UNITED STATES
SECURITIES AND EXCHANGE COMMISSION
Washington, D.C. 20549

 

FORM 6-K

 

REPORT OF FOREIGN PRIVATE ISSUER PURSUANT TO RULE 13a-16 OR 15d-16

UNDER THE SECURITIES EXCHANGE ACT OF 1934

 

For the month of August 2022

 

Commission File Number 1-32135

 

SEABRIDGE GOLD INC.
(Name of Registrant)

 

106 Front Street East, Suite 400, Toronto, Ontario, Canada M5A 1E1
(Address of Principal Executive Office)

 

Indicate by check mark whether the registrant files or will file annual reports under cover of Form 20-F or Form 40-F.

 

Form 20-F Form 40-F

 

Indicate by check mark if the registrant is submitting the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(1):

 

Indicate by check mark if the registrant is submitting the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(7):

 

This Report on Form 6-K includes as Exhibit 99.2 the Company’s technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” with an effective date of August 8, 2022. Exhibit 99.2 shall be deemed to be filed and shall be incorporated by reference into the Company’s Registration Statements on Form S-8 (File No. 333-211331) and Form F-10 (File No. 333-251081)

 

 

 

 

 

 

SIGNATURE

 

Pursuant to the requirements of the Securities Exchange Act of 1934, the registrant has duly caused this report to be signed on its behalf by the undersigned, thereunto duly authorized.

 

  Seabridge Gold Inc.
  (Registrant)
     
Date: August 10, 2022 By: /s/ Chris Reynolds
  Name:  Chris Reynolds
  Title: VP Finance and CFO

 

1

 

   

EXHIBITS

 

Exhibit   Description
99.1   Material Change Report dated August 9, 2022
99.2   Technical Report
23.1   Consent of Henry Kim, P.Geo., Wood Canada Limited.
23.2   Consent of Hassan Ghaffari, P.Eng., Tetra Tech, Inc
23.3   Consent of Jianhui (John) Huang, Ph.D., P.Eng., Tetra Tech, Inc.
23.4   Consent of James H. Gray, P.Eng., Moose Mountain Technical Services
23.5   Consent of Derek Kinakin, M.Sc., P.Geo., P.G., BGC Engineering Inc.
23.6   Consent of Rolf Schmitt, P. Geo., ERM Consultants Canada Ltd.
23.7   Consent of Neil Brazier, P.Eng., WN Brazier Associates Inc.
23.8   Consent of David Willms, P.Eng., Klohn Crippen Berger Ltd.
23.9   Consent of Ross Hammett, Ph.D., P.Eng., WSP Golder

 

 

2

 

 

 

Exhibit 23.1

 

August 10, 2022

 

TO:Seabridge Gold Inc.

United States Securities and Exchange Commission

 

Re:Seabridge Gold Inc. (the “Company”)

Consent of Expert

 

Ladies and Gentlemen:

 

Reference is made to the technical report titled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” dated August 8, 2022 (the “Technical Report”) prepared for the Company.

 

The undersigned hereby consents to:

 

1.the public filing of the Technical Report by the Company;

 

2.any extracts from or a summary of the Technical Report contained in, or incorporated by reference in:

 

a.the Company’s news release dated June 28, 2022,

 

b.the Company’s news release dated August 8, 2022,

 

c.the Company’s Amended and Restated Material Change Report dated August 9, 2022 in respect of the Technical Report,

 

d.the Company’s report on Form 6-K dated August 10, 2022,

 

e.the Company’s Short Form Base Shelf Prospectus dated December 3, 2020 and the Prospectus Supplement dated January 22, 2021 thereto, and

 

f.the Company’s Registration Statements on Form S-8 (File No. 333-211331) and Form F-10, as amended (File No. 333-251081),

 

(collectively the “Disclosure Documents”);

 

3.being named directly or indirectly in the Disclosure Documents; and

 

4.the use of the Technical Report and to extracts from or a summary of the Technical Report in the Disclosure Documents or incorporated by reference therein.

 

The undersigned hereby certifies and confirms that:

 

i.the undersigned has read the Disclosure Documents, including the extracts from or a summary of the Technical Report in the Disclosure Documents, collectively, or incorporated by reference therein;

 

ii.the undersigned has no reason to believe that there are any misrepresentations in the information contained in the Disclosure Documents or incorporated by reference therein that is derived from the Technical Report or that is within the undersigned’s knowledge as a result of the services performed by the undersigned in connection with the Technical Report; and

 

iii.the Disclosure Documents, collectively fairly and accurately represent the information in the sections of the Technical Report for which I am responsible.

 

  Yours Truly,
   
  “Henry H. Kim”
  Henry H. Kim, P.Geo
  Senior Resource Geologist
  Wood Canada Limited

 

 

Exhibit 23.2

 

August 10, 2022

 

TO:Seabridge Gold Inc.

United States Securities and Exchange Commission

 

Re:Seabridge Gold Inc. (the “Company”)

Consent of Expert

 

Ladies and Gentlemen:

 

Reference is made to the technical report titled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” dated August 8, 2022 (the “Technical Report”) prepared for the Company.

 

The undersigned hereby consents to:

 

1.the public filing of the Technical Report by the Company;

 

2.any extracts from or a summary of the Technical Report contained in, or incorporated by reference in:

 

a.the Company’s news release dated June 28, 2022,

 

b.the Company’s news release dated August 8, 2022,

 

c.the Company’s Amended and Restated Material Change Report dated August 9, 2022 in respect of the Technical Report,

 

d.the Company’s report on Form 6-K dated August 10, 2022,

 

e.the Company’s Short Form Base Shelf Prospectus dated December 3, 2020 and the Prospectus Supplement dated January 22, 2021 thereto, and

 

f.the Company’s Registration Statements on Form S-8 (File No. 333-211331) and Form F-10, as amended (File No. 333-251081),

 

(collectively the “Disclosure Documents”);

 

3.being named directly or indirectly in the Disclosure Documents; and

 

4.the use of the Technical Report and to extracts from or a summary of the Technical Report in the Disclosure Documents or incorporated by reference therein.

 

The undersigned hereby certifies and confirms that:

 

i.the undersigned has read the Disclosure Documents, including the extracts from or a summary of the Technical Report in the Disclosure Documents, collectively, or incorporated by reference therein;

 

ii.the undersigned has no reason to believe that there are any misrepresentations in the information contained in the Disclosure Documents or incorporated by reference therein that is derived from the Technical Report or that is within the undersigned’s knowledge as a result of the services performed by the undersigned in connection with the Technical Report; and

 

iii.the Disclosure Documents, collectively fairly and accurately represent the information in the sections of the Technical Report for which I am responsible.

 

  Yours Truly,
   
  “Hassan Ghaffari”
  Hassan Ghaffari, P.Eng
  Director of Metallurgy
  Tetra Tech Inc.

Exhibit 23.3

 

August 10, 2022

 

TO:Seabridge Gold Inc.

United States Securities and Exchange Commission

 

Re:Seabridge Gold Inc. (the “Company”)

Consent of Expert

 

Ladies and Gentlemen:

 

Reference is made to the technical report titled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” dated August 8, 2022 (the “Technical Report”) prepared for the Company.

 

The undersigned hereby consents to:

 

1.the public filing of the Technical Report by the Company;

 

2.any extracts from or a summary of the Technical Report contained in, or incorporated by reference in:

 

a.the Company’s news release dated June 28, 2022,

 

b.the Company’s news release dated August 8, 2022,

 

c.the Company’s Amended and Restated Material Change Report dated August 9, 2022 in respect of the Technical Report,

 

d.the Company’s report on Form 6-K dated August 10, 2022,

 

e.the Company’s Short Form Base Shelf Prospectus dated December 3, 2020 and the Prospectus Supplement dated January 22, 2021 thereto, and

 

f.the Company’s Registration Statements on Form S-8 (File No. 333-211331) and Form F-10, as amended (File No. 333-251081),

 

(collectively the “Disclosure Documents”);

 

3.being named directly or indirectly in the Disclosure Documents; and

 

4.the use of the Technical Report and to extracts from or a summary of the Technical Report in the Disclosure Documents or incorporated by reference therein.

 

The undersigned hereby certifies and confirms that:

 

i.the undersigned has read the Disclosure Documents, including the extracts from or a summary of the Technical Report in the Disclosure Documents, collectively, or incorporated by reference therein;

 

ii.the undersigned has no reason to believe that there are any misrepresentations in the information contained in the Disclosure Documents or incorporated by reference therein that is derived from the Technical Report or that is within the undersigned’s knowledge as a result of the services performed by the undersigned in connection with the Technical Report; and

 

iii.the Disclosure Documents, collectively fairly and accurately represent the information in the sections of the Technical Report for which I am responsible.

 

  Yours Truly,
   
  “Jianhui Huang”
  Jianhui (John) Huang, Ph.D, P.Eng.
  Senior Metallurgist
  Tetra Tech Inc.

 

 

Exhibit 23.4

 

August 10, 2022

 

TO:Seabridge Gold Inc.

United States Securities and Exchange Commission

 

Re:Seabridge Gold Inc. (the “Company”)

Consent of Expert

 

Ladies and Gentlemen:

 

Reference is made to the technical report titled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” dated August 8, 2022 (the “Technical Report”) prepared for the Company.

 

The undersigned hereby consents to:

 

1.the public filing of the Technical Report by the Company;

 

2.any extracts from or a summary of the Technical Report contained in, or incorporated by reference in:

 

a.the Company’s news release dated June 28, 2022,

 

b.the Company’s news release dated August 8, 2022,

 

c.the Company’s Amended and Restated Material Change Report dated August 9, 2022 in respect of the Technical Report,

 

d.the Company’s report on Form 6-K dated August 10, 2022,

 

e.the Company’s Short Form Base Shelf Prospectus dated December 3, 2020 and the Prospectus Supplement dated January 22, 2021 thereto, and

 

f.the Company’s Registration Statements on Form S-8 (File No. 333-211331) and Form F-10, as amended (File No. 333-251081),

 

(collectively the “Disclosure Documents”);

 

3.being named directly or indirectly in the Disclosure Documents; and

 

4.the use of the Technical Report and to extracts from or a summary of the Technical Report in the Disclosure Documents or incorporated by reference therein.

 

The undersigned hereby certifies and confirms that:

 

i.the undersigned has read the Disclosure Documents, including the extracts from or a summary of the Technical Report in the Disclosure Documents, collectively, or incorporated by reference therein;

 

ii.the undersigned has no reason to believe that there are any misrepresentations in the information contained in the Disclosure Documents or incorporated by reference therein that is derived from the Technical Report or that is within the undersigned’s knowledge as a result of the services performed by the undersigned in connection with the Technical Report; and

 

iii.the Disclosure Documents, collectively fairly and accurately represent the information in the sections of the Technical Report for which I am responsible.

 

  Yours Truly,
   
  “James H. Gray”
  James H. Gray, P.Eng
  Principal Mining Engineer
  Moose Mountain Technical Services

 

 

Exhibit 23.5

 

August 10, 2022

 

TO:Seabridge Gold Inc.

United States Securities and Exchange Commission

 

Re:Seabridge Gold Inc. (the “Company”)

Consent of Expert

 

Ladies and Gentlemen:

 

Reference is made to the technical report titled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” dated August 8, 2022 (the “Technical Report”) prepared for the Company.

 

The undersigned hereby consents to:

 

1.the public filing of the Technical Report by the Company;

 

2.any extracts from or a summary of the Technical Report contained in, or incorporated by reference in:

 

a.the Company’s news release dated June 28, 2022,

 

b.the Company’s news release dated August 8, 2022,

 

c.the Company’s Amended and Restated Material Change Report dated August 9, 2022 in respect of the Technical Report,

 

d.the Company’s report on Form 6-K dated August 10, 2022,

 

e.the Company’s Short Form Base Shelf Prospectus dated December 3, 2020 and the Prospectus Supplement dated January 22, 2021 thereto, and

 

f.the Company’s Registration Statements on Form S-8 (File No. 333-211331) and Form F-10, as amended (File No. 333-251081),

 

(collectively the “Disclosure Documents”);

 

3.being named directly or indirectly in the Disclosure Documents; and

 

4.the use of the Technical Report and to extracts from or a summary of the Technical Report in the Disclosure Documents or incorporated by reference therein.

 

The undersigned hereby certifies and confirms that:

 

i.the undersigned has read the Disclosure Documents, including the extracts from or a summary of the Technical Report in the Disclosure Documents, collectively, or incorporated by reference therein;

 

ii.the undersigned has no reason to believe that there are any misrepresentations in the information contained in the Disclosure Documents or incorporated by reference therein that is derived from the Technical Report or that is within the undersigned’s knowledge as a result of the services performed by the undersigned in connection with the Technical Report; and

 

iii.the Disclosure Documents, collectively fairly and accurately represent the information in the sections of the Technical Report for which I am responsible.

 

  Yours Truly,
   
  “Derek Kinakin”
  Derek Kinakin, P.Geo., P.L. Eng., P.G.
  Principal Engineering Geologist
  BCG Engineering Inc

Exhibit 23.6

 

August 10, 2022

 

TO:Seabridge Gold Inc.

United States Securities and Exchange Commission 

 

Re:Seabridge Gold Inc. (the “Company”)

Consent of Expert

 

Ladies and Gentlemen:

 

Reference is made to the technical report titled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” dated August 8, 2022 (the “Technical Report”) prepared for the Company.

 

The undersigned hereby consents to:

 

1.the public filing of the Technical Report by the Company;

 

2.any extracts from or a summary of the Technical Report contained in, or incorporated by reference in:

 

a.the Company’s news release dated June 28, 2022,

 

b.the Company’s news release dated August 8, 2022,

 

c.the Company’s Amended and Restated Material Change Report dated August 9, 2022 in respect of the Technical Report,

 

d.the Company’s report on Form 6-K dated August 10, 2022,

 

e.the Company’s Short Form Base Shelf Prospectus dated December 3, 2020 and the Prospectus Supplement dated January 22, 2021 thereto, and

 

f.the Company’s Registration Statements on Form S-8 (File No. 333-211331) and Form F-10, as amended (File No. 333-251081),

 

(collectively the “Disclosure Documents”);

 

3.being named directly or indirectly in the Disclosure Documents; and

 

4.the use of the Technical Report and to extracts from or a summary of the Technical Report in the Disclosure Documents or incorporated by reference therein.

 

The undersigned hereby certifies and confirms that:

 

i.the undersigned has read the Disclosure Documents, including the extracts from or a summary of the Technical Report in the Disclosure Documents, collectively, or incorporated by reference therein;

 

ii.the undersigned has no reason to believe that there are any misrepresentations in the information contained in the Disclosure Documents or incorporated by reference therein that is derived from the Technical Report or that is within the undersigned’s knowledge as a result of the services performed by the undersigned in connection with the Technical Report; and

 

iii.the Disclosure Documents, collectively fairly and accurately represent the information in the sections of the Technical Report for which I am responsible.

 

  Yours Truly,
   
  “Rolf Schmitt”
  Rolf Schmitt, P.Geo
  Technical Director, Permitting
  ERM Consultants Canada Ltd.

 

Exhibit 23.7

 

August 10, 2022

 

TO:Seabridge Gold Inc.

United States Securities and Exchange Commission

 

Re:Seabridge Gold Inc. (the “Company”)

Consent of Expert

 

Ladies and Gentlemen:

 

Reference is made to the technical report titled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” dated August 8, 2022 (the “Technical Report”) prepared for the Company.

 

The undersigned hereby consents to:

 

1.the public filing of the Technical Report by the Company;

 

2.any extracts from or a summary of the Technical Report contained in, or incorporated by reference in:

 

a.the Company’s news release dated June 28, 2022,

 

b.the Company’s news release dated August 8, 2022,

 

c.the Company’s Amended and Restated Material Change Report dated August 9, 2022 in respect of the Technical Report,

 

d.the Company’s report on Form 6-K dated August 10, 2022,

 

e.the Company’s Short Form Base Shelf Prospectus dated December 3, 2020 and the Prospectus Supplement dated January 22, 2021 thereto, and

 

f.the Company’s Registration Statements on Form S-8 (File No. 333-211331) and Form F-10, as amended (File No. 333-251081),

 

(collectively the “Disclosure Documents”);

 

3.being named directly or indirectly in the Disclosure Documents; and

 

4.the use of the Technical Report and to extracts from or a summary of the Technical Report in the Disclosure Documents or incorporated by reference therein.

 

The undersigned hereby certifies and confirms that:

 

i.the undersigned has read the Disclosure Documents, including the extracts from or a summary of the Technical Report in the Disclosure Documents, collectively, or incorporated by reference therein;

 

ii.the undersigned has no reason to believe that there are any misrepresentations in the information contained in the Disclosure Documents or incorporated by reference therein that is derived from the Technical Report or that is within the undersigned’s knowledge as a result of the services performed by the undersigned in connection with the Technical Report; and

 

iii.the Disclosure Documents, collectively fairly and accurately represent the information in the sections of the Technical Report for which I am responsible.

 

  Yours Truly,
   
  “Neil Brazier”
  Neil Brazier, P.Eng
  Principal
  WN Brazier Associates Inc

 

Exhibit 23.8

  

August 10, 2022

 

TO:Seabridge Gold Inc.

United States Securities and Exchange Commission

 

Re:Seabridge Gold Inc. (the “Company”)

Consent of Expert

 

Ladies and Gentlemen:

 

Reference is made to the technical report titled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” dated August 8, 2022 (the “Technical Report”) prepared for the Company.

 

The undersigned hereby consents to:

 

1.the public filing of the Technical Report by the Company;

 

2.any extracts from or a summary of the Technical Report contained in, or incorporated by reference in:

 

a.the Company’s news release dated June 28, 2022,

 

b.the Company’s news release dated August 8, 2022,

 

c.the Company’s Amended and Restated Material Change Report dated August 9, 2022 in respect of the Technical Report,

 

d.the Company’s report on Form 6-K dated August 10, 2022,

 

e.the Company’s Short Form Base Shelf Prospectus dated December 3, 2020 and the Prospectus Supplement dated January 22, 2021 thereto, and

 

f.the Company’s Registration Statements on Form S-8 (File No. 333-211331) and Form F-10, as amended (File No. 333-251081),

 

(collectively the “Disclosure Documents”);

 

3.being named directly or indirectly in the Disclosure Documents; and

 

4.the use of the Technical Report and to extracts from or a summary of the Technical Report in the Disclosure Documents or incorporated by reference therein.

 

The undersigned hereby certifies and confirms that:

 

i.the undersigned has read the Disclosure Documents, including the extracts from or a summary of the Technical Report in the Disclosure Documents, collectively, or incorporated by reference therein;

 

ii.the undersigned has no reason to believe that there are any misrepresentations in the information contained in the Disclosure Documents or incorporated by reference therein that is derived from the Technical Report or that is within the undersigned’s knowledge as a result of the services performed by the undersigned in connection with the Technical Report; and

 

iii.the Disclosure Documents, collectively fairly and accurately represent the information in the sections of the Technical Report for which I am responsible.

 

  Yours Truly,
   
  “David A. Willms”
  David A. Willms, P.Eng
  Senior Geotechnical Engineer
  Klohn Crippen Berger Ltd.

 

Exhibit 23.9

 

August 10, 2022

 

TO:Seabridge Gold Inc.

United States Securities and Exchange Commission

 

Re:Seabridge Gold Inc. (the "Company")

 

Ladies and Gentlemen:

 

Reference is made to the technical report titled "KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report" dated August 8, 2022 (the "Technical Report") prepared for the Company, of which the undersigned is responsible for preparing or supervising the preparation of a part.

 

The undersigned hereby consents to:

 

1.the public filing of the Technical Report by the Company;

 

2.any extracts from or a summary of parts of the Technical Report for which the undersigned is responsible contained in, or incorporated by reference in:

 

a.the Company's news release dated August 8, 2022 (the “August News Release”),

 

b.the Company’s Amended and Restated Material Change Report dated August 9, 2022 in respect of the Technical Report (the “Amended and Restated Material Change Report”),

 

c.the Company’s report on Form 6-K dated August 10, 2022 (the “Form 6-K”),

 

d.the Company’s Short Form Base Shelf Prospectus dated December 3, 2020 (the “Prospectus”), and

 

e.the Company’s Registration Statements on Form S-8 (File No. 333-211331) and Form F-10, as amended (File No. 333-251081) (collectively the “Registration Statement”),

 

(collectively the "Disclosure Documents"); and

 

3.being named directly or indirectly in the Disclosure Documents; and

 

4.the use of the parts of the Technical Report for which I am responsible in the Amended and Restated Material Change Report, the Form 6-K, the Prospectus and the Registration Statements.

 

The undersigned hereby confirms that:

 

i.the undersigned has read the Disclosure Documents, including the extracts from or a summary of the Technical Report in the Disclosure Documents;

 

ii.the undersigned has no reason to believe that there are any misrepresentations in the information contained in the Amended and Restated Material Change Report, the Form 6-K, the Prospectus or the Registration Statements that is derived from the parts of the Technical Report for which the undersigned is responsible or that is within the undersigned’s knowledge as a result of the services performed by the undersigned in connection with the Technical Report; and

 

iii.the August News Release fairly and accurately represents the information in the sections of the Technical Report for which I am responsible.

 

  Yours Truly,
   
  “Ross D. Hammett”
  Ross D. Hammett, Ph.D., P.Eng
  Senior Engineer & Principal
  WSP Golder Inc.

Exhibit 99.1

 

FORM 51-102F3

Amended and Restated Material Change Report

 

The short form base shelf prospectus dated December 3, 2020 of Seabridge Gold Inc. is amended and supplemented by the contents of this material change report.

 

ITEM 1. NAME AND ADDRESS OF COMPANY

 

Seabridge Gold Inc. (“Seabridge” or the “Company”)
Suite 400 – 106 Front Street East
Toronto, ON M5A 1E1

 

ITEM 2. DATE OF MATERIAL CHANGE

 

June 28, 2022 and August 8, 2022

 

ITEM 3. NEWS RELEASE

 

The news releases announcing the material change referred to in this report were disseminated on June 28, 2022, August 3, 2022 and August 8, 2022 through Newsfile and copies have been filed under Seabridge’s profile on SEDAR.

 

ITEM 4. SUMMARY OF MATERIAL CHANGE

 

On June 28, 2022, the Company announced the results of an updated Preliminary Feasibility Study (the “2022 PFS”) for its 100% owned KSM project located in northern British Columbia, Canada. The 2022 PFS shows a considerably more sustainable and profitable mining operation than its 2016 predecessor, now consisting of an all open pit mine plan that includes the Mitchell, East Mitchell and Sulphurets deposits only. The primary reasons for the improvements in the plan arise from the acquisition of the East Mitchell open pit resource and an expansion to planned mill throughput. The many design improvements over the 2016 PFS include a smaller environmental footprint, reduced waste rock production, reduced greenhouse gas emissions by electrification of the mine haul fleet, a 50% increase in mill throughput, and the elimination of capital-intensive block cave mining.

 

Notable improvements in the Base Case put forward in the 2022 PFS compared to the Base Case in the previous 2016 PFS include:

 

Proven and probable gold reserves increase 22%, from 38.8 million ounces to 47.3 million ounces, due to higher gold grades added from the East Mitchell deposit.

 

Mill throughput expands from 130,000 tonnes per day (“tpd”) to 195,000 tpd

 

Waste to ore strip ratio reduced by 23% to approximately 1.05:1.

 

A 90% increase in average annual gold production, 14% increase in annual copper production, 37% increase in annual silver production, and a 263% increase in annual molybdenum production1

 

Total capital of US$10.5 billion is reduced to US$9.6 billion with increases from inflation and mill expansion being wholly offset by the elimination of block cave mining from the PFS plan.

 

Initial capital increases from US$5.0 billion to US$6.4 billion primarily due to inflation.

 

 

1Figures corrected from news release dated June 28, 2022. The news release dated June 28, 2022 attached hereto has been corrected to reflect a 14% increase in annual copper production, a 37% increase in annual silver production and a 263% increase in annual molybdenum production.

 

 

 

 

A 20 year reduction in mine life from 53 years to 33 years due to the increased mill throughput supplied by higher open pit production.

 

Total after tax net cash flow increases from US$10.0 billion to US$23.9 billion.

 

After tax NPV(5%) increases from US$1.5 billion to US$7.9 billion.

 

After tax IRR increases from 8.0% to 16.1%.

 

Payback period drops from 6.8 years to 3.7 years.

 

The 2022 PFS was prepared by Tetra Tech, Inc., the firm that had also authored the 2016 PFS. The 2022 PFS results released herein propose mining only 25% of the KSM resource inventory and do not include material from the copper-rich Kerr and Iron Cap deposits.

 

The Technical Report that has been prepared presenting the 2022 PFS also includes a preliminary economic assessment (the “PEA”) in respect of the potential to extend production at the KSM Project from its Kerr and the Iron Cap deposits and the results of the PEA were announced on August 3, 2022.

 

On August 8, 2022, the Issuer announced that it has filed a technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” dated August 8, 2022 (the “2022 Technical Report”) summarized in the June 28, 2022 and August 8, 2022 news releases of the Company. The authors of the 2022 Technical Report are as follows:

 

Hassan Ghaffari, P.Eng., Tetra Tech, Inc.
Jianhui (John) Huang, Ph.D., P.Eng., Tetra Tech, Inc.
Henry Kim, P.Geo., Wood Canada Limited
James H. Gray, P.Eng., Moose Mountain Technical Services
Neil Brazier, P.Eng., WN Brazier Associates Inc.
Rolf Schmitt, P. Geo., ERM Consultants Canada Ltd.
David Willms, P.Eng., Klohn Crippen Berger Ltd.
Derek Kinakin, M.Sc., P.Geo., P.G., BGC Engineering Inc.
Ross Hammett, Ph.D., P.Eng., WSP Golder

 

ITEM 5. FULL DESCRIPTION OF MATERIAL CHANGE

 

See attached news releases.

 

ITEM 6. RELIANCE ON SUBSECTION 7.1(2) or (3) OF NATIONAL INSTRUMENT 51-102

 

This report is not being filed on a confidential basis.

 

ITEM 7. OMITTED INFORMATION

 

There are no significant facts required to be disclosed herein which have been omitted.

 

ITEM 8. EXECUTIVE OFFICER

 

  Contact: Rudi Fronk, Chief Executive Officer
  Telephone: (416) 367-9292

 

ITEM 9. DATE OF REPORT

 

August 9, 2022

 

-2-

 

 

CERTIFICATE OF SEABRIDGE GOLD INC.

 

Dated: August 9, 2022

 

The short form prospectus dated December 3, 2020, as amended by this amendment, together with the documents incorporated in the prospectus by reference, will, as of the date of the last supplement to this prospectus relating to the securities offered by the prospectus and the supplement(s), constitute full, true and plain disclosure of all material facts relating to the securities offered by the prospectus and the supplement(s) as required by the securities legislation of each of the provinces of Ontario, Alberta, British Columbia, Manitoba, Saskatchewan, Nova Scotia and in the Yukon territory.

 

“Rudi P. Fronk”   “Christopher J. Reynolds”
Chief Executive Officer, Chairman & Director   Vice President, Finance & Chief Financial Officer
     
On behalf of the Board of Directors
   

 

“John Sabine”   “Clement Pelletier”
Director   Director

 

-3-

 

 

CERTIFICATE OF THE AGENT

 

Dated: August 9, 2022

 

To the best of our knowledge, information and belief, the short form prospectus dated December 3, 2020, as amended by this amendment, together with the documents incorporated in the prospectus by reference, will, as of the date of the last supplement to the prospectus relating to the securities offered by the prospectus and the supplement(s), constitute full, true and plain disclosure of all material facts relating to the securities offered by the prospectus and the supplement(s) as required by the securities legislation of Ontario, Alberta, British Columbia, Manitoba, Saskatchewan, Nova Scotia and in the Yukon territory.

 

Cantor Fitzgerald & Co

 

  “Sage Kelly”
  Name:  Sage Kelly
  Title: Head of Investment Banking

 

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News Release

 

Trading Symbols: TSX: SEA For Immediate Release
NYSE: SA June 28, 2022

 

Seabridge Gold Completes Updated Preliminary Feasibility Study for KSM Project

 

Estimated Annual Gold Output Rises 90% to more than 1 Million Oz for 33 years

 

After-Tax NPV(5%) soars 426% to $7.9B and Payback Period Shrinks 46% to 3.7 Years

 

Base Case Operating Costs Estimated at US$275 Per Oz of Gold Produced

after copper, silver and molybdenum credits

 

Reserves of 47.3 Million Oz Gold, 7.3 Billion Lbs Copper, 160 Million Oz Silver

and 385 Million Lbs molybdenum

 

Base Case Total Cost (Including all Capital, Reclamation and Closure Costs) Estimated at US$601 Per Oz of Gold Produced after copper, silver and molybdenum credits

 

Toronto, Canada – Seabridge Gold announced today the results of an updated Preliminary Feasibility Study (the “2022 PFS”) for its 100% owned KSM project located in northern British Columbia, Canada. The 2022 PFS shows a considerably more sustainable and profitable mining operation than its 2016 predecessor, now consisting of an all open pit mine plan that includes the Mitchell, East Mitchell and Sulphurets deposits only. The primary reasons for the improvements in the plan arise from the acquisition of the East Mitchell open pit resource and an expansion to planned mill throughput. The many design improvements over the 2016 PFS include a smaller environmental footprint, reduced waste rock production, reduced green house gas emissions by electrification of the mine haul fleet, a 50% increase in mill throughput, and the elimination of capital-intensive block cave mining.

 

The 2022 PFS was prepared by Tetra Tech, Inc. (“Tetra Tech”), the firm that had also authored the 2016 PFS. The 2022 PFS results released herein propose mining only 25% of the KSM resource inventory and do not include material from the copper-rich Kerr and Iron Cap deposits. An analysis of a stand-alone development of these deposits will be included as a Preliminary Economic Assessment (“PEA”) forming a separate part of the NI 43-101 Technical Report to be filed within the next 45 days.

 

Seabridge Gold Chairman and CEO Rudi Fronk noted: “We have redesigned KSM for an inflationary environment. The themes for this PFS are capital and energy efficiency. The mine plan is simplified to bring total capital down below 2016 estimates despite inflation by reducing sustaining capital. We have accomplished this by eliminating underground mine development which is deferred to future years. Important steps have also been taken to make the project less dependent on oil, especially diesel fuel, which is an inflationary hot spot and likely to remain so. We have done this by maximizing the use of low cost, green hydroelectric energy.”

 

 

106 Front Street East, Suite 400, Toronto, ON M5A 1E1, Canada

Telephone: 416-367-9292 www.seabridgegold.com

 

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Notable improvements in the Base Case 2022 PFS compared to the Base Case 2016 PFS include:

 

  Proven and probable gold reserves increase 22%, from 38.8 million ounces to 47.3 million ounces (2,292 million tonnes at 0.64 g/t), due to higher gold grades added from the East Mitchell deposit.
     
  Mill throughput expands from 130,000 tonnes per day (“tpd”) to 195,000 tpd
     
  Waste to ore strip ratio reduced by 23% to approximately 1.05:1.
     
  A 90% increase in average annual gold production, 14% increase in annual copper production, 37% increase in annual silver production, and a 263% increase in annual molybdenum production.(1)
     
  Total capital of US$10.5 billion is reduced to US$9.6 billion with increases from inflation and mill expansion being wholly offset by the elimination of block cave mining from the PFS plan.
     
  Initial capital increases from US$5.0 billion to US$6.4 billion primarily due to inflation.
     
  A 20 year reduction in mine life from 53 years to 33 years due to the increased mill throughput supplied by higher open pit production.
     
  Total after tax net cash flow increases from US$10.0 billion to US$23.9 billion.
     
  After tax NPV(5%) increases from US$1.5 billion to US$7.9 billion.
     
  After tax IRR increases from 8.0% to 16.1%.
     
  Payback period drops from 6.8 years to 3.7 years.

 

The 2022 PFS envisages an open pit mine operation that is scheduled to operate for 33 years. Ore delivery to the mill is increased from an initial 130,000 metric tpd to 195,000 tpd in Year 3. Over the entire 33-year mine life, ore will be fed to a flotation and gold extraction mill. The flotation plant will produce a gold/copper/silver concentrate for transport by truck to a nearby seaport at Stewart, B.C. for shipment to Pacific Rim smelters. Metallurgical projections supported by extensive metallurgical testing project a copper concentrate with an average copper grade of 24% and a high gold (64 g/t) and silver (177g/t) content, making it readily saleable. A separate molybdenum concentrate and gold-silver doré will be produced at the KSM processing facility.

 

Mineral Resources

 

The 2022 PFS uses previously disclosed resource estimates that are based on US$1,300 per ounce gold, US$3.00 per pound copper, US$20.00 per ounce silver and US$9.70 per pound molybdenum. In addition, the resources are constrained by conceptual mining shapes.

 

Measured and Indicated Mineral Resources at KSM are estimated at 5.4 billion tonnes grading 0.51 grams per tonne gold, 0.16% copper, 2.4 grams per tonne silver, and 63 ppm molybdenum (88.3 million ounces of gold, 19.4 billion pounds of copper, 414 million ounces of silver, and 742 million pounds of molybdenum). An additional 5.7 billion tonnes are estimated in the Inferred Mineral Resource category grading 0.36 grams per tonne gold, 0.28% copper, 2.2 grams per tonne silver, and 33 ppm molybdenum (65.6 million ounces of gold, 35.1 billion pounds of copper and 406 million ounces of silver, and 415 million pounds of molybdenum). A detailed table of KSM’s mineral resources can be found at the end of this news release

 

Mineral Reserves

 

Updated Mineral Reserves for the project are based on open pit mining of the Mitchell, East Mitchell and Sulphurets deposits. Waste to ore cut-offs were determined using a net smelter return (“NSR”) for each block in the model. NSR is calculated using prices and process recoveries for each metal accounting for all off-site losses, transportation, smelting and refining charges. Metal prices of US$1,300 per ounce gold, US$3.00 per pound copper, US$20 per ounce silver and US$9.70 per pound molybdenum and a foreign exchange rate of 0.79 US dollar per Canadian dollar have been used in the NSR calculations.

 

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Lerchs-Grossman (“LG”) pit shell optimizations were used to define open pit mine pit limits in the 2022 PFS. Open pit designed phases use updated geotechnical design criteria based on recent site investigation programs. Mineral Reserves have been estimated using the updated pit designs. The open pit minimum NSR cut-off grade is varied between Cdn$ 11/t to Cdn$25/t and considers the estimated process operating cost of Cdn$10/t. Process operating costs include plant processing (including crushing/conveying costs where applicable), G&A, surface service, tailings, and water treatment costs. A premium cut-off grade of Cdn$25/t is used until the end of Year 5 to maximize the NPV and minimize the time to payback of initial capital.

 

Mineral Reserves for the KSM project are stated as follows:

 

KSM Proven and Probable Mineral Reserves as of May 26, 2022

 

  Ore
(Mt)
Diluted Grades Contained Metal
Au
(g/t)
Cu
(%)
Ag
(g/t)
Mo
(ppm)
Au
(Moz)
Cu
(Mlb)
Ag
(Moz)
Mo
(Mlb)
Proven Mitchell 483 0.74 0.20 3.3 49 11.5 2,161 51 53
East Mitchell 814 0.69 0.11 1.8 91 18.1 2,043 47 163
Sulphurets 0 0.00 0.00 0.0 0 0.0 0 0 0
Total Proven 1,297 0.71 0.15 2.4 75 29.6 4,203 98 215
Probable Mitchell 452 0.59 0.15 2.5 74 8.6 1,458 36 74
East Mitchell 392 0.46 0.09 1.7 84 5.8 784 21 73
Sulphurets 151 0.68 0.26 1.0 70 3.3 874 5 23
Total Probable 995 0.55 0.14 1.9 77 17.7 3,116 62 170
Proven + Probable Mitchell 935 0.67 0.18 2.9 61 20.1 3,619 87 126
East Mitchell 1,206 0.62 0.11 1.8 89 23.9 2,826 68 236
Sulphurets 151 0.68 0.26 1.0 70 3.3 874 5 23
Total Proven + Probable 2,292 0.64 0.14 2.2 76 47.3 7,320 160 385

 

Notes:

 

1.The Mineral Reserve estimates were reviewed by Jim Gray, P.Eng. (who is also the independent Qualified Person for these Mineral Reserve estimates), reported using the 2014 CIM Definition Standards and 2019 CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines, and have an effective date of May 26, 2022.
2.Mineral Reserves are based on the 2022 PFS all open pit Life of Mine plan.
3.Mineral Reserves are mined tonnes and grade, the reference point is the mill feed at the primary crusher and includes consideration for operational modifying factors.
4.Mineral Reserves are reported at NSR cut-off grades that vary between of $11/t and $25/t using the following assumptions: metal prices of US$1300/oz Au, US$3.00/lb Cu, US$20/oz Ag, and US$ 9.70/lb Mo at a currency exchange rate of 0.79 US$ per CAD$; Copper concentrate terms are 96% payable Cu; 97.8% payable Au; 90% payable Ag, molybdenum concentrate terms are 99% payable. Offsite costs (smelting, refining, transport, and insurance) are Cdn$281 per tonne of copper concentrate and Cdn$5527 per tonne of molybdenum concentrate; doré terms are $2/oz offsite costs (refining, transport and insurance), 99.8% Au payable, and 90% Ag payable; metallurgical recovery projections vary depending on metallurgical domain and metal grades and are based on metallurgical test work.
5.The NSR cut-off grade is varied from Cdn11/t to Cdn25/t and covers the estimated process operating cost of $10/t for ore processing, G&A, surface service, tailings , and water treatment costs.
6.Mineral Reserves account for mining loss and dilution.
7.Mineral Reserves are a subset of the mineral resource.
8.Numbers have been rounded as required by reporting guidelines.

 

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Production

 

The open pit only mine production plan using ultra class mining starts in the higher grade Mitchell pit. Production from the high grade upper East Mitchell zone is introduced in Year 3. Waste mined from the Sulphurets, East Mitchell and Mitchell pit is placed in the Mitchell rock storage facility (RSF) until Mitchell pit is mined out by Year 25. Final waste from East Mitchell is backfilled into the mined out Mitchell pit from Year 25 onward along with some waste rehandled from the Mitchell RSF.

 

The updated mine plan reduces overall footprint by not using the McTagg RSF as required in the 2016 PFS and by utilizing mined out pits for backfilling waste rock.

 

Autonomous mine operations where applicable and an integrated remote operations centre reduce on-site personnel.

 

Electrification of the haul truck fleet with trolley assist reduces carbon emissions and overall mine energy costs by replacing diesel with low cost energy from electricity.

 

Mill feed ramps up to 130,000 tonnes per day by Year 2 followed by a 50% increase to 195,000 tonnes per day from Year 3 onwards. Average annual mill feed throughput for the 33 years of mine life is estimated at 69.5 million tonnes.

 

At Mitchell, a near-surface higher grade gold zone crops out allowing for gold production in the first seven years that is substantially above the mine life average. The mine plan is specifically designed for mining highest gold grade first to facilitate a quick capital investment payback. The project’s post-tax payback period is approximately 3.7 years for the Base Case or 11% of mine life. Metal production for the first seven years, compared to life of mine average production, is estimated as follows:

 

Average Annual Metal Production

 

  Years 1-7
Average
Life of Mine
Average

Average Grades:

Gold (grams per tonne)

Copper (%)

Silver (grams per tonne)

Molybdenum (parts per million)

 

0.89

0.21

3.0

52

 

0.64

0.14

2.2

76

Annual Production:

Gold (ounces)

Copper (pounds)

Silver (ounces)

Molybdenum (pounds)

 

1,413,000

251 million

3.8 million

2.1 million

 

1,027,000

178 million

3.0 million

4.2 million

 

Note: Annual production shows total metal contained in copper concentrate, doré, and molybdenum concentrate.

 

Capital Costs

 

Initial capital cost (including contingency of US$ 949 million) is estimated at US$6.4 billion, approximately 28% higher than the initial capital estimate in the 2016 PFS primarily due to inflation experienced over the past two years. Initial capital assumes certain early works (e.g. roads and power infrastructure) are being completed ahead of a major project construction decision as a part of the ongoing KSM substantial start activities.

 

Sustaining capital over the 33 year mine life is estimated at US$3.2 billion, a reduction of US$2.3 billion from the 2016 PFS, and is dominated by mill throughput expansion and mine fleet ramp up in Year 1 and 2, and tailings sustaining capital mid way through the mine life.

 

In addition to sustaining capital, a further US$1,273million has been charged against the project for a sinking fund during the production period to pay for estimated water treatment obligations which continue after closure and for physical reclamation and post closure maintenance after mining operations have ceased. 

 

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Initial capital and sustaining capital estimates are summarized as follows:

 

Capital Costs (US$ million)

 

  Initial Sustaining Total
US$ M US$ M US$ M
Direct Costs      
  Mine 1,420 1,766 3,187
  Process 2,003 309 2,312
  Tailings Management Facility 513 630 1,143
  Environmental 15 8 23
  On-site Infrastructure 39 - 39
  Off-site Infrastructure 76 11 87
  Power Supply/Energy Recovery 121 46 167
Total Direct Capital 4,188 2,770 6,958
  Indirect cost 1,090 97 1,188
  Owner’s cost 204 - 204
  Contingency 949 343 1,293
Total Capital 6,432 3,210 9,642

 

Operating Costs

 

Average mine, process and G&A operating costs over the project’s life (including waste mining and on-site power credits, excluding off-site shipping and smelting costs) are estimated at US$11.36 per tonne milled (before base metal credits). Estimated unit operating costs decreased 8% from the 2016 PFS primarily due to the change from combined open pit and block cave mining to open pit only mining, a 50% increase in mill throughput capacity, and technology improvements including automation and electrification of the mine fleet. A breakdown of estimated unit operating costs is as follows:

 

LOM Average Unit Operating Costs (US$ Per Tonne Milled)

 

  Mining 3.31
  Process 6.31
  G&A + Site Services 1.06
  Tailings Storage/Handling 0.11
  Water Management/Treatment 0.50
  Energy Recovery -0.07
  Provincial Sales Tax 0.13
  Total Operating Costs 11.36
   

 

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Economic Analysis

 

A Base Case economic evaluation was undertaken incorporating historical three-year trailing averages for metal prices as of June 20, 2022. This approach is consistent with the 2016 PFS Base Case. Two alternate cases are also presented: (i) an Alternate Case that incorporates lower metal prices than used in the Base Case to demonstrate the project’s sensitivity to lower prices; and (ii) a Recent Spot Case incorporating recent spot prices for gold, copper, silver and the US$/Cdn$ exchange rate. The pre-tax and post-tax estimated economic results in U.S. dollars for all three cases as well as the 2016 PFS Base Case are as follows:

 

Projected Economic Results (US$)

 

  2016 PFS
Base Case
2022 PFS
Base Case
2022 PFS Recent Spot Case 2022 PFS
Alternate
Case
Metal Prices:        
    Gold ($/ounce) 1,230 1,742 1,850 1,500
Copper ($/pound) 2.75 3.53 4.25 3.00

Silver ($/ounce)

Molybdenum ($/lb)

17.75

8.49

21.90

18.00

22.00

18.00

20.00

18.00

US$/Cdn$ Exchange Rate: 0.80 0.77 0.77 0.77
Cost Summary:        
Operating Costs Per Ounce of Gold Produced (years 1 to 7) $119 $35 -$83 $118
Operating Costs Per Ounce of Gold Produced (life of mine) $277 $275 $164 $351
Total Cost Per Ounce of Gold Produced (inclusive of all capital and closure) $673 $601 $490 $677

Initial Capital (billions)

Sustaining Capital (billions)

$5.0

$5.5

$6.4

$3.2

$6.4

$3.2

$6.4

$3.2

Unit Operating Cost (US$/tonne) $12.36 $11.36 $11.36 $11.36
Pre-Tax Results:        
    Net Cash Flow (billions) $15.9  $38.6  $46.1  $27.9
    NPV @ 5% Discount Rate (billions) $3.3  $13.5  $16.4  $9.2
    Internal Rate of Return 10.4% 20.1% 22.4% 16.5%
    Payback Period (years) 6.0  3.4  3.1  4.1
Post-Tax Results:        
    Net Cash Flow (billions) $10.0  $23.9  $28.6  $17.1
    NPV @ 5% Discount Rate (billions) $1.5  $7.9  $9.8  $5.2
    Internal Rate of Return 8.0% 16.1% 18.0% 13.1%
    Payback Period (years) 6.8 3.7 3.4 4.3

 

Note:

 

1.Operating and total cost per ounce of gold are after copper, silver and molybdenum credits.
2.Total cost per ounce includes all start-up capital, sustaining capital and reclamation/closure costs.
3.Results include consideration of Royalties and Impact Benefit Agreements
4.The post-tax results include the B.C. Mineral Tax and provincial and federal corporate taxes.

 

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The NI 43-101 Technical Report will include sensitivity analyses illustrating the impact on project economics from positive and negative changes to metal prices, capital costs and operating costs.

 

National Instrument 43-101 Disclosure The updated KSM PFS was prepared by Tetra Tech, and incorporates the work of a number of industry-leading consulting firms. These firms and their Qualified Persons (as defined under National Instrument 43-101) are independent of Seabridge and have reviewed and approved this news release. The principal consultants who contributed to the 2022 PFS, and their Qualified Persons are listed below along with their areas of responsibility:

 

Tetra Tech, under the direction of Hassan Ghaffari P.Eng (surface infrastructure, capital estimate and financial analysis), John Huang P.Eng. (metallurgical testing review, permanent water treatment, mineral process design and operating cost estimation for process, G&A and site services, and overall report preparation)

 

Wood Canada Limited under the direction of Henry Kim P.Geo. (Mineral Resources)

 

Moose Mountain Technical Services under the direction of Jim Gray P.Eng. (open pit Mineral Reserves, open pit mining operations, mine capital and mine operating costs, MTT and rail ore conveyance design, tunnel capital costs)

 

W.N. Brazier Associates Inc. under the direction of W.N. Brazier P.Eng. (Electrical power supply, energy recovery plants)

 

ERM (Environmental Resources Management) under the direction of Rolf Schmitt P.Geo. (environment and permitting)

 

Klohn Crippen Berger Ltd. Under the direction of David Willms P.Eng (design of surface water diversions, diversion tunnels, tailing management facility, water treatment dam and RSF and tunnel geotechnical)

 

BGC Engineering Inc. under the direction of Derek Kinakin P.Geo., P.L.Eng. (rock mechanics, geohazards and mining pit slopes)

 

Seabridge holds a 100% interest in several North American gold projects. Seabridge’s assets include the KSM and Iskut projects located near Stewart, British Columbia, Canada, the Courageous Lake project located in Canada’s Northwest Territories, the Snowstorm project in the Getchell Gold Belt of Northern Nevada and the 3 Aces project set in the Yukon Territory. For a full breakdown of Seabridge’s Mineral Reserves and Mineral Resources by category please visit the Company’s website at http://www.seabridgegold.com.

 

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Neither the Toronto Stock Exchange, New York Stock Exchange, nor their Regulation Services Providers accepts responsibility for the adequacy or accuracy of this release.

 

All reserve and resource estimates reported by the Corporation were estimated in accordance with the Canadian National Instrument 43-101 and the Canadian Institute of Mining, Metallurgy and Petroleum (“CIM”) Definition Standards. The U.S. Securities and Exchange Commission (“SEC”) now recognizes estimates of “measured mineral resources,” “indicated mineral resources” and “inferred mineral resources” and uses new definitions of “proven mineral reserves” and “probable mineral reserves” that are substantially similar to the corresponding CIM Definition Standards. However, the CIM Definition Standards differ from the requirements applicable to US domestic issuers. US investors are cautioned not to assume that any “measured mineral resources,” “indicated mineral resources,” or “inferred mineral resources” that the Issuer reports are or will be economically or legally mineable. Further, “inferred mineral resources” are that part of a mineral resource for which quantity and grade are estimated on the basis of limited geologic evidence and sampling. Mineral resources which are not mineral reserves do not have demonstrated economic viability.

 

This document contains “forward-looking information” within the meaning of Canadian securities legislation and “forward-looking statements” within the meaning of the United States Private Securities Litigation Reform Act of 1995. This information and these statements, referred to herein as “forward-looking statements” are made as of the date of this document. Forward-looking statements relate to future events or future performance and reflect current estimates, predictions, expectations or beliefs regarding future events and include, but are not limited to, statements with respect to: (i) the estimated amount and grade of mineral reserves and mineral resources; (ii) estimates of the capital costs of constructing mine facilities and bringing a mine into production, of operating the mine, of sustaining capital and the duration of financing payback periods; (iii) the estimated amount of future production, both ore processed and metal recovered; (iv) estimates of operating costs, life of mine costs, net cash flow, net present value (NPV) and economic returns from an operating mine; and (v) the completion of a Preliminary Economic Assessment and its timing. Any statements that express or involve discussions with respect to predictions, expectations, beliefs, plans, projections, objectives or future events or performance (often, but not always, using words or phrases such as “expects”, “anticipates”, “plans”, “projects”, “estimates”, “envisages”, “assumes”, “intends”, “strategy”, “goals”, “objectives” or variations thereof or stating that certain actions, events or results “may”, “could”, “would”, “might” or “will” be taken, occur or be achieved, or the negative of any of these terms and similar expressions) are not statements of historical fact and may be forward-looking statements.

 

All forward-looking statements are based on Seabridge’s or its consultants’ current beliefs as well as various assumptions made by them and information currently available to them. The most significant assumptions are set forth above, but other these assumptions include: (i) the presence of and continuity of metals at the Project at estimated grades; (ii) the geotechnical and metallurgical characteristics of rock conforming to sampled results; (iii) the quantities of water and the quality of the water that must be diverted or treated during mining operations; (iv) the capacities and durability of various machinery and equipment; (v) the availability of personnel, machinery, equipment and hydro-electric power at estimated prices and within the estimated delivery times; (v) currency exchange rates; (vi) metals sales prices; (vii) appropriate discount rates applied to the cash flows in the economic analysis; (viii) tax rates and royalty rates applicable to the proposed mining operation; (ix) the availability of acceptable financing under assumed structure and costs; (ix) anticipated mining losses and dilution; (x) metallurgical performance; (xi) reasonable contingency requirements; (xii) success in realizing proposed operations; (xiii) receipt of permits and other regulatory approvals on acceptable terms; and (xiv) the successful conclusion of consultation with impacted indigenous groups. Although management considers these assumptions to be reasonable based on information currently available to it, they may prove to be incorrect. Many forward-looking statements are made assuming the correctness of other forward-looking statements, such as statements of net present value and internal rates of return, which are based on most of the other forward-looking statements and assumptions herein. The cost information is also prepared using current values, but the time for incurring the costs will be in the future and it is assumed costs (and metals prices) will remain stable over the relevant period.

 

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By their very nature, forward-looking statements involve inherent risks and uncertainties, both general and specific, and risks exist that estimates, forecasts, projections and other forward-looking statements will not be achieved or that assumptions do not reflect future experience. We caution readers not to place undue reliance on these forward-looking statements as a number of important factors could cause the actual outcomes to differ materially from the beliefs, plans, objectives, expectations, anticipations, estimates assumptions and intentions expressed in such forward-looking statements. These risk factors may be generally stated as the risk that the assumptions and estimates expressed above do not occur as forecast, but specifically include, without limitation: risks relating to variations in the mineral content within the material identified as mineral reserves or mineral resources from that predicted; variations in rates of recovery and extraction; the geotechnical characteristics of the rock mined or through which infrastructure is built differing from that predicted, the quantity of water that will need to be diverted or treated during mining operations being different from what is expected to be encountered during mining operations or post closure, or the rate of flow of the water being different; developments in world metals markets; risks relating to fluctuations in the Canadian dollar relative to the US dollar; increases in the estimated capital and operating costs or unanticipated costs; difficulties attracting the necessary work force; unavailability of hydro-electric power and risks relating to the costs of other energy sources; increases in financing costs or adverse changes to the terms of available financing, if any; tax rates or royalties being greater than assumed; changes in development or mining plans due to changes in logistical, technical or other factors; changes in project parameters as plans continue to be refined; risks relating to receipt of regulatory approvals or the conclusion of successful consultation with impacted indigenous groups; changes in regulations applying to the development, operation, and closure of mining operations from what currently exists; the effects of competition in the markets in which Seabridge operates; operational and infrastructure risks and the additional risks described in Seabridge’s Annual Information Form filed with SEDAR in Canada (available at www.sedar.com ) for the year ended December 31, 2021 and in the Corporation’s Annual Report Form 40-F filed with the U.S. Securities and Exchange Commission on EDGAR (available at www.sec.gov/edgar.shtml). Seabridge cautions that the foregoing list of factors that may affect future results is not exhaustive.

 

When relying on our forward-looking statements to make decisions with respect to Seabridge, investors and others should carefully consider the foregoing factors and other uncertainties and potential events. Seabridge does not undertake to update any forward-looking statement, whether written or oral, that may be made from time to time by Seabridge or on our behalf, except as required by law.

 

  ON BEHALF OF THE BOARD
  “Rudi Fronk”
  Chairman and C.E.O.

 

For further information please contact:
Rudi P. Fronk, Chairman and C.E.O.
Tel: (416) 367-9292 • Fax: (416) 367-2711
Email: info@seabridgegold.com

 

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KSM Project Mineral Resources (Inclusive of Mineral Reserves as stated above)

 

Measured Resources

 

Project Cut Off Grade (g/t) Tonnes (000) Gold Copper Silver Molybdenum
Grade (g/t) Ounces (000) Grade (%) Pounds (millions) Grade (g/t) Ounces (000) Grade (ppm) Pounds (millions)
KSM: NSR:                  
  Mitchell $10.75 691,700 0.68 15,124 0.19 2,876 3.3 72,831 52 79
  East Mitchell $11.25 1,012,800 0.65 21,098 0.11 2,514 1.8 59,233 89 198
KSM Total   1,704,500 0.66 36,222 0.14 5,390 2.4 132,064 74 277

 

Indicated Resources

 

Project Cut Off
Grade
(g/t)
Tonnes (000) Gold Copper Silver Molybdenum
Grade (g/t) Ounces (000) Grade (%) Pounds (millions) Grade (g/t) Ounces (000) Grade (ppm) Pounds (millions)
KSM:

$10.75-$11.25 NSR Pits

 

$16 NSR

UG

                 
  Mitchell 1,667,000 0.48 25,935 0.14 5,120 2.8 149,160 66 241
  East Mitchell 746,200 0.42 10,080 0.08 1,390 1.7 41,814 79 130
  Sulphurets 446,000 0.55 7,887 0.21 2,064 1.0 14,339 53 52
  Kerr 374,000 0.22 2,660 0.41 3,405 1.1 13,744 5 4
  Iron Cap 423,000 0.41 5,576 0.22 2,051 4.6 62,559 41 38
KSM Total 3,656,200 0.44 52,138 0.17 14,030 2.4 281,616 58 465

 

Measured plus Indicated Resources

 

Project Cut Off
Grade
(g/t)
Tonnes (000) Gold Copper Silver Molybdenum
Grade (g/t) Ounces (000) Grade (%) Pounds (millions) Grade (g/t) Ounces (000) Grade (ppm) Pounds (millions)
KSM:

$10.75-$11.25 NSR Pits

 

$16 NSR

UG

                 
  Mitchell 2,358,700 0.54 41,059 0.15 7,996 2.9 221,991 62 320
  East Mitchell 1,759,000 0.55 31,178 0.10 3,904 1.8 101,047 85 328
  Sulphurets 446,000 0.55 7,887 0.21 2,064 1.0 14,339 53 52
  Kerr 370,000 0.22 2,660 0.41 3,405 1.1 13,744 5 4
  Iron Cap 423,000 0.41 5,576 0.22 2,051 4.6 62,559 41 38
KSM Total 5,356,700 0.51 88,360 0.16 19,420 2.4 413,680 63 742

 

-14-

 

 

Inferred Resources

 

Project Cut Off Grade (g/t) Tonnes
(000)
Gold Copper Silver Molybdenum
Grade (g/t) Ounces (000) Grade (%) Pounds (millions) Grade (g/t) Ounces (000) Grade (ppm) Pounds (millions)
KSM:

$10.75 NSR Pits

 

$16 NSR

UG

                 
  Mitchell 1,282,600 0.29 11,819 0.14 3,832 2.5 102,228 47 133
  East Mitchell 281,100 0.37 3,372 0.07 403 2.3 21,112 61 38
  Sulphurets 223,000 0.44 3,155 0.13 639 1.3 9,320 30 15
  Kerr 1,999,000 0.31 19,823 0.40 17,720 1.8 114,431 23 103
  Iron Cap 1,899,000 0.45 27,474 0.30 12,556 2.6 158,741 30 126
KSM Total 5,684,700 0.36 65,643 0.28 35,150 2.2 405,832 33 415

 

Note:

 

1.The effective date for the Mineral Resource Estimate for Mitchell and East Mitchell is March 31, 2022, and for Kerr, Sulphurets and Iron Cap is December 31, 2019.
2.The Mineral Resource estimates have been reviewed and approved by Henry Kim P.Geo., an independent Qualified Person. Mr. Kim verified the databases supporting the mineral resource estimates and conducted a personal inspection of the property and reviewed drill core from a range of representative drill holes at site and at the core storage facilities in Stewart, B.C. with Seabridge geology staff.
3.Mineral Resources were prepared in accordance with CIM Definition Standards for Mineral Resources and Mineral Reserves (May 10, 2014) and CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines (Nov 29, 2019).
4.Mineral Resources were constrained within minable shapes depending on their mining methods.
5.Mineral Resources are reported inclusive of those Mineral Resources that were converted to Mineral Reserves. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.
6.Following metal prices were used to determine Mineral Resources: US$1300/oz Au, US$3/lb Cu, US$20/oz Ag, and US$ 9.7/lb Mo.
7.For other key assumption parameters, methods used for: Mitchell and East Mitchell, see news release “Seabridge Gold Reports Updated Mineral Resource Estimates for Mitchell and East Mitchell Deposits” dated April 14, 2022; Kerr, Sulphuret, and Iron Cap, see “KSM (KERR-SULPHURETS-MITCHELL) PREFEASIBILITY STUDY UPDATE, NI 43-101 TECHNICAL REPORT” dated April 30, 2020.
8.Numbers may not add due to rounding.

 

Note: United States investors are cautioned that the requirements and terminology of NI 43-101 may differ from the requirements of the SEC, including Regulation SK-1300. Accordingly, the Issuer’s disclosures regarding mineralization may not be comparable to similar information disclosed by companies subject to the SEC’s mining disclosure standards. Mineral Resources are reported inclusive of Mineral Reserves. Mineral Resources which are not Mineral Reserves do not have demonstrated economic viability. It is reasonably expected that the majority of Inferred Mineral Resources could be upgraded to Indicated Mineral Resources with continued exploration.

 

-15-

 

 

 

News Release

 

Trading Symbols: TSX: SEA For Immediate Release
NYSE: SA August 3, 2022

 

New KSM Preliminary Economic Assessment (“PEA”) Sees Additional

Copper-Rich Block Cave Opportunity

PEA Based on Kerr and Iron Cap Deposits Not Included in Recently Updated PFS

39 Year PEA Mill Feed of 1.7 Bt Contains 16 B lb Copper and 23.2 Moz Gold

 

Base Case Operating Costs Estimated at US$0.38 Per Lb of Copper Produced after gold, silver and molybdenum credits

 

Base Case Total Cost (Including all Capital) Estimated at US$1.44 Per Lb of Copper Produced after gold, silver and molybdenum credits


After Tax NPV5% US$5.8 B, After Tax IRR 18.9%, After Tax Payback in 6.2 Years


Toronto, Canada – Seabridge Gold announced today the results of a Preliminary Economic Assessment Study (the “2022 PEA”) for a potential copper-rich underground mine at its 100% owned KSM project located in northern British Columbia, Canada. The 2022 PEA is a stand-alone mine plan that has been undertaken to evaluate a potential future expansion of the KSM mine to the copper rich Iron Cap and Kerr deposits after the 2022 Preliminary Feasibility Study (“PFS”) mine plan has been completed. The 2022 PEA is primarily an underground block cave mining operation supplemented with a small open pit and is planned to operate for 39 years with a peak mill feed production of 170,000 t/d, demonstrating that KSM has multigenerational long-life mining project potential with flexibility to vary metal output.

 

The 2022 PFS plan disclosed on 28 June 2022 is an open pit only plan with a 33 year mine life limited to the Mitchell, East Mitchell, and Sulphurets deposits. None of the mineral resources incorporated into the 2022 PEA mine plan have been used in the 2022 PFS mine plan. For the news release announcing the 2022 PFS results please click here.

 

Seabridge Gold Chairman and CEO Rudi Fronk noted: “KSM is really an entire district hosting a nest of potentially economic porphyry deposits with different characteristics. In our updated PFS we focused on the gold-rich deposits because of their faster payback and the relative simplicity of an open-pit only operation. However, we are very mindful that a deep deficit in mined copper is projected to be on the horizon as the world electrifies and moves towards a net zero carbon future. We therefore wanted to highlight KSM’s potential to contributute to addressing this need more fully than the mine plan contained in our updated Preliminary Feasibilty Study. We think this opportunity will be attractive to a prospective partner.”

 

The 2022 PEA envisages an underground focussed mine plan starting with the development of an Iron Cap block cave mine supplemented with a small open pit at Kerr. Development of a Kerr block cave mine begins when Iron Cap development tapers off. Kerr block cave mill feed starts 6 years after the start of Iron Cap mill feed. Mill feed delivery to the process plant is ramped up to 170,000 tpd by Year 12. Over the entire 39-year mine life, mill feed will be delivered to a flotation concentration mill circuit. The flotation plant will produce a gold/copper/silver concentrate and separate molybdenum concentrate for transport by truck to a nearby seaport at Stewart, B.C.

 

-16-

 

 

Mineral Resources

 

The 2022 PFS and the 2022 PEA uses previously disclosed resource estimates that are based on US$1,300 per ounce gold, US$3.00 per pound copper, US$20.00 per ounce silver and US$9.70 per pound molybdenum. In addition, the resources are constrained by conceptual mining shapes.

 

Measured and Indicated Mineral Resources at KSM are estimated at 5.4 billion tonnes grading 0.51 grams per tonne gold, 0.16% copper, 2.4 grams per tonne silver, and 63 ppm molybdenum (88.4 million ounces of gold, 19.4 billion pounds of copper, 414 million ounces of silver, and 742 million pounds of molybdenum). An additional 5.7 billion tonnes are estimated in the Inferred Mineral Resource category grading 0.36 grams per tonne gold, 0.28% copper, 2.2 grams per tonne silver, and 33 ppm molybdenum (65.6 million ounces of gold, 35.2 billion pounds of copper and 406 million ounces of silver, and 415 million pounds of molybdenum). A detailed table of KSM’s mineral resources can be found at the end of this news release.

 

2022 PEA Mine Design

 

Kerr open pit has been designed to supplement block cave mill feed during the ramp up of the PEA block cave production.

 

Waste to mill feed cut-offs are determined using a Net Smelter Return (“NSR”) for each block in the model. The pit delineated resources for the 2022 PEA use an NSR cut-off of Cdn$10.75/t. NSR is calculated using prices and process recoveries for each metal accounting for all off-site losses, transportation, smelting and refining charges. Metal prices of US$1,200 per ounce gold, US$2.70 per pound copper, and US$17.50 per ounce silver and a foreign exchange rate of US$ 0.83 per Cdn$1.00 are used in the NSR calculations.

 

The underground block caving mine designs for Iron Cap and Kerr are based on modeling using GEOVIA’s Footprint Finder (FF) software. The ramp-up and maximum yearly mine production rates are established based on the rate at which the drawpoints are constructed and the assumptions are conservatively less than the demonstrated maximum industry rate and the initial and maximum production rates at which individual drawpoints can be mucked. The values chosen for these inputs are based on industry averages adjusted to suit the anticipated conditions.

 

The Iron Cap block cave mine includes an estimated development duration of 4 years, a production ramp-up period of 6 years, steady state production at 32.9 million tonnes per year for 17 years, and then a production ramp-down period of 6 years. The Iron Cap block cave is located adjacent to the Mitchell-Treaty Tunnels (“MTT”), the transportation conduit between mine and mill.

 

The Iron Cap mine is designed as a partially electrified mine with partial automation where battery electric vehicles replace diesel production loaders on the extraction level and trains replace trucks on the haulage level. The height of draw averages around 500m, ranging from 200m on the west limit that is developed early in the mine life to 750m on the east edge of the design that is developed late in the mine life.

 

The Kerr block cave has an estimated development duration of five years, a production ramp-up period of 5 years, and steady state production at 29.2 million tonnes per year for 20 total years with a seven year production dip to as low as 15.0 million tonnes during the transition from the first to second lift.

 

The Kerr block cave has been designed as a conventionally developed and operated block cave mine leaving additional upside for improvement by electrification.

 

The mining NSR shut-off is Cdn$20 per tonne for the Iron Cap block cave and Cdn$18 per tonne for the Kerr block cave. The mill feed contained in the mine plan for the 2022 PEA including dilution and mining losses are stated as follows.

 

-17-

 

 

Mill Feed from the PEA Mine Plan

 

Zone Mining Method Classification Tonnes (millions) Average Grades Contained Metal

Gold

(g/t)

Copper

(%)

Silver

(g/t)

Gold

M oz’s

Copper

M lbs

Silver

M oz’s

Iron Cap Block Cave M+I 58 0.62 0.28 3.2 1.1 354 5.9
Inferred 685 0.58 0.36 3.0 12.7 5,424 65.4
Kerr Open Pit M+I 117 0.26 0.51 1.4 1.0 1,315 5
Inferred 7 0.74 0.09 1.5 0.2 14 0
Block Cave M+I 48 0.25 0.53 1.3 0.4 557 2.0
Inferred 777 0.31 0.49 1.7 7.8 8,339 43.6
Total Mill Feed Mined M+I 223 0.35 0.45 1.8 2.5 2,226 13
Inferred 1,469 0.44 0.43 2.3 20.7 13,777 109

 

Note: The 2022 PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the results of the 2022 PEA will be realized. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

 

Production

 

The 2022 PEA assumes that the 2022 PFS plan has been completed. Open pit mining equipment will be relocated to the Kerr deposit to begin prestripping while the Iron Cap block cave is being developed. Year 1 of the 2022 PEA mine life coincides with the first year of mill feed from the Iron Cap deposit. Mill feed from Kerr block cave begins in Year 7. The 2022 PEA production plan produces 14.3 Billion pounds of copper, 14.3 Million ounces of gold, 68.2 million ounces of silver, and 13.8 million pounds of molybdenum from 1.7 Billion tonnes of mill feed over a 39 year mine life. The production schedule is shown in the graph below.

 

2022 PEA Mill Feed Production Schedule

 

 

-18-

 

 

Average annual production is summarized estimated as follows:

 

Average Annual Metal Production

 

  Life of Mine
Average

Average Grades:

Gold (grams per tonne)

Copper (%)

Silver (grams per tonne)

Molybdenum (parts per million)

 

0.43

0.43

2.2

24

Average Annual Production:

Gold (ounces)

Copper (pounds)

Silver (ounces)

Molybdenum (pounds)

 

368,000

366 million

1.8 million

0.4 million

 

Note: Annual production shows total metal contained in copper concentrate, doré, and molybdenum concentrate.

 

Tailing management is envisioned as a combination of technically viable storage approaches that will be refined in future studies to comprise appropriate and responsible solutions depending on best selected locations and available technology.

 

Capital Costs

 

Initial capital cost for the 2022 PEA is estimated at US$1.5 billion with sustaining capital over the 39 year mine life estimated at US$12.8 billion dominated by block cave development capital. Initial capital includes all capital until the first year of mill feed (Year 1). Capital estimates are summarized as follows:

 

2022 PEA Capital Costs (US$ million)

 

  Initial Sustaining Total
US$ M US$ M US$ M
Direct Costs      
  Mine 828 6,678 7,506
  Process 0 651 651
  Tailings Management Facility 74 664 738
  On-site Infrastructure 26 573 599
  Power Supply/Energy Recovery 0 112 112
Total Direct Capital 927 8,678 9,606
  Indirect cost 253 1249 1,502
  Contingency 320 2824 3,145
Total Capital 1,500 12,752 14,252

 

Note: Numbers may not add due to rounding

 

-19-

 

 

Operating Costs

 

Average mine, process and G&A operating costs over the project’s life (including waste mining and on-site power credits, excluding off-site shipping and smelting costs) are estimated at US$11.98 per tonne milled (before base metal credits). A breakdown of estimated unit operating costs is as follows:

 

2022 PEA LOM Average Unit Operating Costs (US$ Per Tonne Milled)

 

  Mining 4.99
  Process 4.31
  G&A + Site Services 1.89
  Tailings Storage/Handling 0.15
  Water Management/Treatment 0.68
  Energy Recovery -0.09
  Provincial Sales Tax 0.05
  Total Operating Costs 11.98

 

Economic Analysis

 

A Base Case economic evaluation was undertaken incorporating historical three-year trailing averages for gold, copper and silver metal prices of as of June 20, 2022. This approach is used because it is consistent with the 2022 PFS Base Case. Molybdenum price is based on a recent study for a primary molybdenum project. Two alternate cases are also presented: (i) an Alternate Case that incorporates lower metal prices than used in the Base Case to demonstrate the project’s sensitivity to lower prices; and, (ii) a Recent Spot Case incorporating recent spot prices for gold, copper, silver and the US$/Cdn$ exchange rate. The pre-tax and post-tax estimated economic results in U.S. dollars for all three are as follows:

 

2022 PEA Projected Economic Results (US$)

 

  2022 PEA Base Case 2022 PEA Alternate Case 2022 PEA Recent Spot Case
Metal Prices:      
Gold ($/ounce) 1,742 1,500 1,850
Copper ($/pound) 3.53 3.00 4.25

Silver ($/ounce)

Molybdenum ($/lb)

21.90

18.00

20.00

18.00

22.00

18.00

US$/Cdn$ Exchange Rate: 0.77 0.77 0.77
Cost Summary:      
Operating Costs Per Pound of Copper Produced (life of mine) $0.38 $0.59 $0.32
Total Cost Per Pound of Copper Produced (inclusive of all capital) $1.44 $1.64 $1.38
Pre-Tax Results:      
Net Cash Flow (billions)  $29.8  $19.4  $40.9
NPV @ 5% Discount Rate (billions)  $9.7  $5.8  $13.9
Internal Rate of Return 24.0% 17.4% 30.4%
Payback Period (years)  4.7  7.5  3.9
Post-Tax Results:      
Net Cash Flow (billions)  $18.5  $11.9  $25.6
NPV @ 5% Discount Rate (billions)  $5.8  $3.3  $8.4
Internal Rate of Return 18.9% 13.5% 24.0%
Payback Period (years) 6.2 8.7 4.4

 

Note:

 

5.The 2022 PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the results of the 2022 PEA will be realized. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.
6.Results include consideration of Royalties and Impact Benefit Agreements.
7.Operating and total cost per pound of copper produced are after gold, silver and molybdenum credits.
8.The post-tax results include the B.C. Mineral Tax and provincial and federal corporate taxes.
9.Cash flows are discounted to the start of the 2022 PEA development.
10.Payback years are measured from the first year of mill feed.

 

-20-

 

 

The NI 43-101 Technical Report will include sensitivity analyses illustrating the impact on project economics from positive and negative changes to metal prices, capital costs and operating costs.

 

National Instrument 43-101 Disclosure The 2022 KSM PEA was prepared by Tetra Tech, and incorporates the work of a number of industry-leading consulting firms. These firms and their Qualified Persons (as defined under National Instrument 43-101) are independent of Seabridge and have reviewed and approved this news release. The principal consultants who contributed to the 2022 PEA, and their Qualified Persons are listed below along with their areas of responsibility:

 

Tetra Tech, under the direction of Hassan Ghaffari P.Eng (surface infrastructure, capital estimate and financial analysis), John Huang P.Eng. (metallurgical testing review, permanent water treatment, mineral process design and operating cost estimation for process, G&A and site services, and overall report preparation)

 

Wood Canada Limited under the direction of Henry Kim P.Geo. (Mineral Resources)

 

WSP Golder, under the Direction of Ross Hammett P.Eng (Block Cave mining)

 

Moose Mountain Technical Services under the direction of Jim Gray P.Eng. (open pit mining, MTT and rail mill feed conveyance design, tunnel capital costs)

 

W.N. Brazier Associates Inc. under the direction of W.N. Brazier P.Eng. (Electrical power supply, energy recovery plants)

 

ERM (Environmental Resources Management) under the direction of Rolf Schmitt P.Geo. (environment and permitting)

 

Klohn Crippen Berger Ltd. Under the direction of David Willms P.Eng (design of surface water diversions, diversion tunnels, tailing management facility, water treatment dam and RSF and tunnel geotechnical)

 

Seabridge holds a 100% interest in several North American gold projects. Seabridge’s assets include the KSM and Iskut projects located in Northwest British Columbia, Canada’s “Golden Triangle”, the Courageous Lake project located in Canada’s Northwest Territories, the Snowstorm project in the Getchell Gold Belt of Northern Nevada and the 3 Aces project set in the Yukon Territory. For a full breakdown of Seabridge’s Mineral Reserves and Mineral Resources by category please visit the Company’s website at http://www.seabridgegold.com.

 

Neither the Toronto Stock Exchange, New York Stock Exchange, nor their Regulation Services Providers accepts responsibility for the adequacy or accuracy of this release.

 

All mineral reserve and resource estimates reported by the Corporation were estimated in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (“CIM”) Definition Standards. The U.S. Securities and Exchange Commission (“SEC”) now recognizes estimates of “measured mineral resources,” “indicated mineral resources” and “inferred mineral resources” and uses new definitions of “proven mineral reserves” and “probable mineral reserves” that are substantially similar to the corresponding CIM Definition Standards. However, the CIM Definition Standards differ from the requirements applicable to US domestic issuers. US investors are cautioned not to assume that any “measured mineral resources,” “indicated mineral resources,” or “inferred mineral resources” that the Issuer reports. Further, “inferred mineral resources” are that part of a mineral resource for which quantity and grade are estimated on the basis of limited geologic evidence and sampling. Mineral resources which are not mineral reserves do not have demonstrated economic viability.

 

-21-

 

 

This document contains “forward-looking information” within the meaning of Canadian securities legislation and “forward-looking statements” within the meaning of the United States Private Securities Litigation Reform Act of 1995. This information and these statements, referred to herein as “forward-looking statements” are made as of the date of this document. Forward-looking statements relate to future events or future performance and reflect current estimates, predictions, expectations or beliefs regarding future events and include, but are not limited to, statements with respect to: (i) the estimated amount and grade of mineral resources; (ii) estimates of the capital costs and timing of constructing the facilities for the potential mine expansion and bringing the additional mine areas into production, of operating such mine, of sustaining capital and the duration of capital payback periods; (iii) the estimated amount of future production, both ore processed and metal recovered; (iv) estimates of operating costs, life of mine costs, net cash flow, net present value (NPV) and economic returns from an operating mine; (v) estimates of block cave ramp-up, production and ramp-down rates; (vi) the assumptions on which the various estimates are made are reasonable; and (vii) projections of a future deep deficit in mined copper. Any statements that express or involve discussions with respect to predictions, expectations, beliefs, plans, projections, objectives or future events or performance (often, but not always, using words or phrases such as “expects”, “anticipates”, “plans”, “projects”, “estimates”, “envisages”, “assumes”, “intends”, “strategy”, “goals”, “objectives” or variations thereof or stating that certain actions, events or results “may”, “could”, “would”, “might” or “will” be taken, occur or be achieved, or the negative of any of these terms and similar expressions) are not statements of historical fact and may be forward-looking statements.

 

All forward-looking statements are based on Seabridge’s or its consultants’ current beliefs as well as various assumptions made by them and information currently available to them. The most significant assumptions are set forth above, but these assumptions include: (i) the presence of and continuity of metals at the Property at estimated grades; (ii) the geotechnical and metallurgical characteristics of rock conforming to sampled results; (iii) the quantities of water and the quality of the water that must be diverted or treated during mining operations; (iv) the capacities and durability of various machinery and equipment; (v) the availability of personnel, machinery, equipment and hydro-electric power at estimated prices and within the estimated delivery times; (v) currency exchange rates; (vi) metals sales prices; (vii) appropriate discount rates applied to the cash flows in the economic analysis; (viii) tax rates and royalty rates applicable to the proposed mining operation; (ix) the availability of acceptable financing under assumed structure and costs; (ix) anticipated mining losses and dilution; (x) metallurgical performance; (xi) reasonable contingency requirements; (xii) success in realizing proposed construction and operations timelines; (xiii) receipt of permits and other regulatory approvals on acceptable terms; and (xiv) the successful conclusion of consultation with impacted indigenous groups. Although management considers these assumptions to be reasonable based on information currently available to it, they may prove to be incorrect. Many forward-looking statements are made assuming the correctness of other forward-looking statements, such as statements of net present value and internal rates of return, which are based on most of the other forward-looking statements and assumptions herein. The cost information is also prepared using current values, but the time for incurring the costs will be in the future and it is assumed costs (and metals prices) will remain stable over the relevant period.

 

By their very nature, forward-looking statements involve inherent risks and uncertainties, both general and specific, and risks exist that estimates, forecasts, projections and other forward-looking statements will not be achieved or that assumptions do not reflect future experience. We caution readers not to place undue reliance on these forward-looking statements as a number of important factors could cause the actual outcomes to differ materially from the beliefs, plans, objectives, expectations, anticipations, estimates assumptions and intentions expressed in such forward-looking statements. These risk factors may be generally stated as the risk that the assumptions and estimates expressed above do not occur as forecast, but specifically include, without limitation: risks relating to variations in the mineral content within the material identified as mineral reserves or mineral resources from that predicted; variations in rates of recovery and extraction; the geotechnical characteristics of the rock mined or through which infrastructure is built differing from that predicted, the quantity of water that will need to be diverted or treated during mining operations being different from what is expected to be encountered during mining operations or post closure, or the rate of flow of the water being different; developments in world metals markets; risks relating to fluctuations in the Canadian dollar relative to the US dollar; increases in the estimated capital and operating costs or unanticipated costs; difficulties attracting the necessary work force; unavailability of hydro-electric power and risks relating to the costs of other energy sources; increases in financing costs or adverse changes to the terms of available financing, if any; tax rates or royalties being greater than assumed; changes in development or mining plans due to changes in logistical, technical or other factors; changes in project parameters as plans continue to be refined; risks relating to receipt of regulatory approvals or the conclusion of successful consultation with impacted indigenous groups; changes in regulations applying to the development, operation, and closure of mining operations from what currently exists; the effects of competition in the markets in which Seabridge operates; operational and infrastructure risks and the additional risks described in Seabridge’s Annual Information Form filed with SEDAR in Canada (available at www.sedar.com ) for the year ended December 31, 2021 and in the Corporation’s Annual Report Form 40-F filed with the U.S. Securities and Exchange Commission on EDGAR (available at www.sec.gov/edgar.shtml). Seabridge cautions that the foregoing list of factors that may affect future results is not exhaustive.

 

When relying on our forward-looking statements to make decisions with respect to Seabridge, investors and others should carefully consider the foregoing factors and other uncertainties and potential events. Seabridge does not undertake to update any forward-looking statement, whether written or oral, that may be made from time to time by Seabridge or on our behalf, except as required by law.

 

  ON BEHALF OF THE BOARD
  “Rudi Fronk”
  Chairman and C.E.O.

 

For further information please contact:
Rudi P. Fronk, Chairman and C.E.O.
Tel: (416) 367-9292 • Fax: (416) 367-2711
Email: info@seabridgegold.com

 

-22-

 

 

KSM Project Mineral Resources (Inclusive of Mineral Reserves as stated above)

 

Measured Resources

 

Project Cut Off ($/t) Tonnes (000) Gold Copper Silver Molybdenum
Grade (g/t) Ounces (millions) Grade (%) Pounds (millions) Grade (g/t) Ounces (millions) Grade (ppm) Pounds (millions)
KSM: NSR:                  
  Mitchell $10.75 692,000 0.68 15.1 0.19 2,876 3.3 72.8 52 79
  East Mitchell $11.25 1,013,000 0.65 21.0 0.11 2,514 1.8 59.2 89 198
KSM Total   1,705,000 0.66 36.2 0.14 5,390 2.4 132.0 74 277

 

Indicated Resources

 

Project Cut Off ($/t) Tonnes (000) Gold Copper Silver Molybdenum
Grade (g/t) Ounces (millions) Grade (%) Pounds (millions) Grade (g/t) Ounces (millions) Grade (ppm) Pounds (millions)
KSM:

$10.75-$11.25 NSR Pits

 

$16 NSR

UG

                 
  Mitchell 1,667,000 0.48 25.9 0.14 5,120 2.8 149.2 66 241
  East Mitchell 746,000 0.42 10.0 0.08 1,390 1.7 41.8 79 130
  Sulphurets 446,000 0.55 7.9 0.21 2,064 1.0 14.3 53 52
  Kerr 374,000 0.22 2.7 0.41 3,405 1.1 13.7 5 4
  Iron Cap 423,000 0.41 5.6 0.22 2,051 4.6 62.6 41 38
KSM Total 3,656,000 0.44 52.1 0.17 14,030 2.4 281.6 58 465

 

Measured plus Indicated Resources

 

Project Cut Off ($/t) Tonnes (000) Gold Copper Silver Molybdenum
Grade (g/t) Ounces (millions) Grade (%) Pounds (millions) Grade (g/t) Ounces (millions) Grade (ppm) Pounds (millions)
KSM:

$10.75-$11.25 NSR Pits

 

$16 NSR

UG

                 
  Mitchell 2,359,000 0.54 41.1 0.15 7,996 2.9 222.0 62 320
  East Mitchell 1,759,000 0.55 31.2 0.10 3,904 1.8 101.0 85 328
  Sulphurets 446,000 0.55 7.9 0.21 2,064 1.0 14.3 53 52
  Kerr 370,000 0.22 2.7 0.41 3,405 1.1 13.7 5 4
  Iron Cap 423,000 0.41 5.6 0.22 2,051 4.6 62.6 41 38
KSM Total 5,357,000 0.51 88.4 0.16 19,420 2.4 413.7 63 742

 

-23-

 

 

Inferred Resources

 

Project Cut Off ($/t) Tonnes (000) Gold Copper Silver Molybdenum
Grade (g/t) Ounces (millions) Grade (%) Pounds (millions) Grade (g/t) Ounces (millions) Grade (ppm) Pounds (millions)
KSM:

$10.75 NSR Pits

 

$16 NSR

UG

                 
  Mitchell 1,283,000 0.29 11.8 0.14 3,832 2.5 102.2 47 133
  East Mitchell 281,000 0.37 3.3 0.07 403 2.3 21.1 61 38
  Sulphurets 223,000 0.44 3.2 0.13 639 1.3 9.3 30 15
  Kerr 1,999,000 0.31 19.8 0.40 17,720 1.8 114.4 23 103
  Iron Cap 1,899,000 0.45 27.5 0.30 12,556 2.6 158.7 30 126
KSM Total 5,685,000 0.36 65.6 0.28 35,150 2.2 405.8 33 415

 

Note:

 

9.The effective date for the Mineral Resource Estimate for Mitchell and East Mitchell is March 31, 2022, and for Kerr, Sulphurets and Iron Cap is December 31, 2019.
10.The Mineral Resource estimates have been reviewed and approved by Henry Kim P.Geo., an independent Qualified Person. Mr. Kim verified the databases supporting the mineral resource estimates and conducted a personal inspection of the property and reviewed drill core from a range of representative drill holes at site and at the core storage facilities in Stewart, B.C. with Seabridge geology staff.
11.Mineral Resources were prepared in accordance with CIM Definition Standards for Mineral Resources and Mineral Reserves (May 10, 2014) and CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines (Nov 29, 2019).
12.Mineral Resources were constrained within mineable shapes depending on their mining methods.
13.Mineral Resources are reported inclusive of those Mineral Resources that were converted to Mineral Reserves. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.
14.Following metal prices were used to determine Mineral Resources: US$1300/oz Au, US$3/lb Cu, US$20/oz Ag, and US$ 9.7/lb Mo.
15.For other key assumption parameters, methods used for: Mitchell and East Mitchell, see news release “Seabridge Gold Reports Updated Mineral Resource Estimates for Mitchell and East Mitchell Deposits” dated April 14, 2022; Kerr, Sulphuret, and Iron Cap, see “KSM (KERR-SULPHURETS-MITCHELL) PREFEASIBILITY STUDY UPDATE, NI 43-101 TECHNICAL REPORT” dated April 30, 2020.
16.Numbers may not add due to rounding.

 

Note: United States investors are cautioned that the requirements and terminology of NI 43-101 may differ from the requirements of the SEC, including Regulation SK-1300. Accordingly, the Issuer’s disclosures regarding mineralization may not be comparable to similar information disclosed by companies subject to the SEC’s mining disclosure standards. Mineral Resources are reported inclusive of Mineral Reserves. Mineral Resources which are not Mineral Reserves do not have demonstrated economic viability. It is reasonably expected that the majority of Inferred Mineral Resources could be upgraded to Indicated Mineral Resources with continued exploration.

 

-24-

 

 

 

News Release

 

Trading Symbols: TSX: SEA   For Immediate Release
  NYSE: SA     August 8, 2022

 

Seabridge Gold Files KSM Technical Report

Report Contains PFS with 33 Year Open Pit Mine Plan and

A 39 Year PEA Capturing Higher Grade Copper Deposits Not Included in PFS

 

Toronto, Canada – Seabridge Gold announced today that it has filed a National Instrument 43-101 technical report at www.sedar.com for its 100%-owned KSM project located in northern British Columbia, Canada entitled “KSM Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” with an effective date of August 8, 2022 (the “Technical Report”).

 

Chairman and CEO Rudi Fronk commented that “for the first time we have provided a complete picture of KSM’s potential for multi-generational gold, copper, silver and molybdenum production. This Technical Report enables potential partners to evaluate all the major opportunities and challenges of the project as defined by industry-leading consultants in their respective fields of expertise. The PFS lays out a development option which is simplified and limited to open pit production with a rapid payback thanks in part to a very low strip ratio for such a large mine. Average annual gold production is estimated at more than 1 million ounces per year. The PEA defines a robust copper-rich opportunity from block caving. Both have excellent economics at current metal prices and share innovative environmental enhancements including substantial electrification.”

 

2022 PFS Highlights

 

The 2022 Preliminary Feasibility Study (“2022 PFS”) shows a considerably more sustainable and profitable mining operation than its 2016 predecessor, now consisting of an all open pit mine plan that includes the Mitchell, East Mitchell and Sulphurets deposits only. The primary reasons for the improvements in the plan compared to the 2016 study arise from the acquisition of the East Mitchell open pit resource and an expansion to planned mill throughput. The many design improvements over the 2016 PFS include a smaller environmental footprint, reduced waste rock production, reduced green house gas emissions by partial electrification of the mine haul fleet, a 50% increase in mill throughput, and the elimination of capital-intensive block cave mining.

 

The 2022 PFS has redesigned KSM for an inflationary environment and include capital and energy efficiency. The mine plan was simplified to bring total capital down below the 2016 estimates despite inflation by eliminating sustaining capital associated with block cave development. Important steps were also taken to make the project less dependent on oil, especially diesel fuel, which is an inflationary hot spot and likely to remain so. We have done this by maximizing the use of low cost, green hydroelectric energy.

 

Highlights of the 2022 PFS include:

 

Proven and Probable Mineral Reserves of 47.3 million ounces of Gold, 7.3 billion pounds of copper, 160 million ounces of silver and 385 million pounds of molybdenum

 

Average annual metal production over the 33 year mine life of 1.03 million ounces of gold, 178 million pounds of copper, 3.0 million ounces of silver and 4.2 million pounds of molybdenum

 

Base case operating costs estimated at US$275 per ounce of gold produced after copper, silver and molybdenum credits

 

Base case total cost (Including initial and sustaining capital, reclamation and closure costs) estimated at US$601 per ounce of gold produced after copper, silver and molybdenum credits

 

After-Tax NPV(5%) estimated at US$7.9B, after tax IRR at 16.1% and after tax payback of 3.7 Years

 

-25-

 

 

The 2022 PFS envisages an open pit mine operation that is scheduled to operate for 33 years. Ore delivery to the mill is increased from an initial 130,000 metric tonnes per day (“tpd”) to 195,000 tpd in Year 3. Over the entire 33-year mine life, ore will be fed to a flotation and gold extraction mill. The flotation plant will produce a gold/copper/silver concentrate for transport by truck to a nearby seaport at Stewart, B.C. for shipment to Pacific Rim smelters. Metallurgical projections supported by extensive metallurgical testing project a copper concentrate with an average copper grade of 24% and a high gold (64 g/t) and silver (177g/t) content, making it readily saleable. A separate molybdenum concentrate and gold-silver doré will be produced at the KSM processing facility. Base case metal prices used in the 2022 PFS were US$1,742 per ounce of gold, US$3.53 per pound of copper, US$21.90 per ounce of silver and US$18.00 per pound of molybdenum.

 

For further details on the 2022 PFS please see here.

 

2022 PEA Highlights

 

The 2022 Preliminary Economic Assessment (“2022 PEA”) is a stand-alone mine plan that has been undertaken to evaluate a potential future expansion of the KSM mine to the copper rich Iron Cap and Kerr deposits after the 33-year 2022 PFS mine plan has been completed. The 2022 PEA is primarily an underground block cave mining operation supplemented with a small open pit and is planned to operate for 39 years with a peak mill feed production of 170,000 tpd, demonstrating that KSM has multigenerational long-life mining project potential with flexibility to vary metal output.

 

Highlights of the 2022 PEA using the same base case metal prices as the 2022 PFS include:

 

Based entirely on KSM’s Kerr and Iron Cap deposits that are not included in the 2022 PFS

 

39-year mill feed of 1.7 billion tonnes containing 16 billion pounds of Copper, 23.2 million ounces of gold and 122 million ounces of silver

 

Average annual production over the 39-year mine life of 368,000 ounces of gold, 366 million pounds of copper, 1.8 million ounces of silver and 0.4 million pounds of molybdenum

 

Base case operating costs estimated at US$0.38 per pound of copper produced after gold, silver and molybdenum credits

 

Base case total cost (including all capital) estimated at US$1.44 per pound of copper produced after gold, silver and molybdenum credits

 

After tax NPV5% estimated at US$5.8 billion, after tax IRR at 18.9%, and after-tax payback at 6.2 Years

 

The 2022 PEA envisages an underground focused mine plan starting with the development of an Iron Cap block cave mine supplemented with a small open pit at Kerr. Development of a Kerr block cave would commence after the production ramp-up period of Iron Cap is completed. Kerr block cave mill feed starts 6 years after the start of Iron Cap mill feed. Mill feed delivery to the process plant is ramped up to 170,000 tpd by Year 12. Over the entire 39-year mine life, mill feed will be delivered to a conventional flotation concentrator. The flotation plant will produce a gold/copper/silver concentrate and separate molybdenum concentrate for transport by truck to a nearby seaport at Stewart, B.C.

 

For further details on the 2022 PEA please see here.

 

Note: The 2022 PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the results of the 2022 PEA will be realized. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

 

-26-

 

 

National Instrument 43-101 Disclosure

 

The new Technical Report incorporates the work of a number of industry-leading consulting firms. These firms and their Qualified Persons (as defined under National Instrument 43-101) are independent of Seabridge and have reviewed and approved this news release. The authors of the Technical Report are listed below along with the names of their employers:

 

Tetra Tech, under the direction of Hassan Ghaffari P.Eng (surface infrastructure, capital estimate and financial analysis), John Huang P.Eng. (metallurgical testing review, permanent water treatment, mineral process design and operating cost estimation for process, G&A and site services, and overall report preparation)

 

Wood, under the direction of Henry Kim P.Geo. (Mineral Resources)

 

Moose Mountain Technical Services under the direction of Jim Gray P.Eng. (open pit Mineral Reserves, open pit mining operations, mine capital and mine operating costs, MTT and rail ore conveyance design, tunnel capital costs)

 

W.N. Brazier Associates Inc. under the direction of W.N. Brazier P.Eng. (Electrical power supply, energy recovery plants)

 

ERM (Environmental Resources Management) under the direction of Rolf Schmitt P.Geo. (environment and permitting)

 

Klohn Crippen Berger Ltd. Under the direction of David Willms P.Eng (design of surface water diversions, diversion tunnels, tailing management facility, water treatment dam and RSF and tunnel geotechnical)

 

BGC Engineering Inc. under the direction of Derek Kinakin P.Geo., P.L.Eng. (rock mechanics, geohazards and mining pit slopes)

 

WSP Golder, under the Direction of Ross Hammett P.Eng (Block Cave mining)

 

Seabridge holds a 100% interest in several North American gold projects. Seabridge’s assets include the KSM and Iskut projects located in Northwest British Columbia, Canada’s “Golden Triangle”, the Courageous Lake project located in Canada’s Northwest Territories, the Snowstorm project in the Getchell Gold Belt of Northern Nevada and the 3 Aces project set in the Yukon Territory. For a full breakdown of Seabridge’s Mineral Reserves and Mineral Resources by category please visit the Company’s website at http://www.seabridgegold.com.

 

Neither the Toronto Stock Exchange, New York Stock Exchange, nor their Regulation Services Providers accepts responsibility for the adequacy or accuracy of this release.

 

All mineral reserve and resource estimates reported by the Corporation were estimated in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (“CIM”) Estimation of Mineral Resources and Mineral Reserve Best Practice Guidelines (2019) and the CIM Definition Standards. The U.S. Securities and Exchange Commission (“SEC”) now recognizes estimates of “measured mineral resources,” “indicated mineral resources” and “inferred mineral resources” and uses new definitions of “proven mineral reserves” and “probable mineral reserves” that are substantially similar to the corresponding CIM Definition Standards. However, the CIM Definition Standards differ from the requirements applicable to US domestic issuers. US investors are cautioned not to assume that any “measured mineral resources,” “indicated mineral resources,” or “inferred mineral resources”. Further, “inferred mineral resources” are that part of a mineral resource for which quantity and grade are estimated on the basis of limited geologic evidence and sampling. Mineral resources which are not mineral reserves do not have demonstrated economic viability.

 

This document contains “forward-looking information” within the meaning of Canadian securities legislation and “forward-looking statements” within the meaning of the United States Private Securities Litigation Reform Act of 1995. This information and these statements, referred to herein as “forward-looking statements” are made as of the date of this document. Forward-looking statements relate to future events or future performance and reflect current estimates, predictions, expectations or beliefs regarding future events and include, but are not limited to, statements with respect to: (i) the estimated amount and grade of mineral resources; (ii) estimates of the capital costs and timing of constructing the facilities for the potential mine expansion and bringing the additional mine areas into production, of operating such mine, of sustaining capital and the duration of capital payback periods; (iii) the estimated amount of future production, both ore processed and metal recovered; (iv) estimates of operating costs, life of mine costs, net cash flow, net present value (NPV) and economic returns from an operating mine; (v) estimates of block cave ramp-up, production and ramp-down rates; (vi) the assumptions on which the various estimates are made are reasonable; and (vii) projections of a future deep deficit in mined copper. Any statements that express or involve discussions with respect to predictions, expectations, beliefs, plans, projections, objectives or future events or performance (often, but not always, using words or phrases such as “expects”, “anticipates”, “plans”, “projects”, “estimates”, “envisages”, “assumes”, “intends”, “strategy”, “goals”, “objectives” or variations thereof or stating that certain actions, events or results “may”, “could”, “would”, “might” or “will” be taken, occur or be achieved, or the negative of any of these terms and similar expressions) are not statements of historical fact and may be forward-looking statements.

 

-27-

 

 

All forward-looking statements are based on Seabridge’s or its consultants’ current beliefs as well as various assumptions made by them and information currently available to them. The most significant assumptions are set forth above, but these assumptions include: (i) the presence of and continuity of metals at the Property at estimated grades; (ii) the geotechnical and metallurgical characteristics of rock conforming to sampled results; (iii) the quantities of water and the quality of the water that must be diverted or treated during mining operations; (iv) the capacities and durability of various machinery and equipment; (v) the availability of personnel, machinery, equipment and hydro-electric power at estimated prices and within the estimated delivery times; (v) currency exchange rates; (vi) metals sales prices; (vii) appropriate discount rates applied to the cash flows in the economic analysis; (viii) tax rates and royalty rates applicable to the proposed mining operation; (ix) the availability of acceptable financing under assumed structure and costs; (ix) anticipated mining losses and dilution; (x) metallurgical performance; (xi) reasonable contingency requirements; (xii) success in realizing proposed construction and operations timelines; (xiii) receipt of permits and other regulatory approvals on acceptable terms; and (xiv) the successful conclusion of consultation with impacted indigenous groups. Although management considers these assumptions to be reasonable based on information currently available to it, they may prove to be incorrect. Many forward-looking statements are made assuming the correctness of other forward-looking statements, such as statements of net present value and internal rates of return, which are based on most of the other forward-looking statements and assumptions herein. The cost information is also prepared using current values, but the time for incurring the costs will be in the future and it is assumed costs (and metals prices) will remain stable over the relevant period.

 

By their very nature, forward-looking statements involve inherent risks and uncertainties, both general and specific, and risks exist that estimates, forecasts, projections and other forward-looking statements will not be achieved or that assumptions do not reflect future experience. We caution readers not to place undue reliance on these forward-looking statements as a number of important factors could cause the actual outcomes to differ materially from the beliefs, plans, objectives, expectations, anticipations, estimates assumptions and intentions expressed in such forward-looking statements. These risk factors may be generally stated as the risk that the assumptions and estimates expressed above do not occur as forecast, but specifically include, without limitation: risks relating to variations in the mineral content within the material identified as mineral reserves or mineral resources from that predicted; variations in rates of recovery and extraction; the geotechnical characteristics of the rock mined or through which infrastructure is built differing from that predicted, the quantity of water that will need to be diverted or treated during mining operations being different from what is expected to be encountered during mining operations or post closure, or the rate of flow of the water being different; developments in world metals markets; risks relating to fluctuations in the Canadian dollar relative to the US dollar; increases in the estimated capital and operating costs or unanticipated costs; difficulties attracting the necessary work force; unavailability of hydro-electric power and risks relating to the costs of other energy sources; increases in financing costs or adverse changes to the terms of available financing, if any; tax rates or royalties being greater than assumed; changes in development or mining plans due to changes in logistical, technical or other factors; changes in project parameters as plans continue to be refined; risks relating to receipt of regulatory approvals or the conclusion of successful consultation with impacted indigenous groups; changes in regulations applying to the development, operation, and closure of mining operations from what currently exists; the effects of competition in the markets in which Seabridge operates; operational and infrastructure risks and the additional risks described in Seabridge’s Annual Information Form filed with SEDAR in Canada (available at www.sedar.com ) for the year ended December 31, 2021 and in the Corporation’s Annual Report Form 40-F filed with the U.S. Securities and Exchange Commission on EDGAR (available at www.sec.gov/edgar.shtml). Seabridge cautions that the foregoing list of factors that may affect future results is not exhaustive.

 

When relying on our forward-looking statements to make decisions with respect to Seabridge, investors and others should carefully consider the foregoing factors and other uncertainties and potential events. Seabridge does not undertake to update any forward-looking statement, whether written or oral, that may be made from time to time by Seabridge or on our behalf, except as required by law.

 

  ON BEHALF OF THE BOARD
  “Rudi Fronk”
  Chairman and C.E.O.

 

For further information please contact:
Rudi P. Fronk, Chairman and C.E.O.
Tel: (416) 367-9292 • Fax: (416) 367-2711
Email: 
info@seabridgegold.com

 

 

-28-

 

 

 

 

Exhibit 99.2

 

Prepared for:

 

Seabridge Gold Inc.

 

 

 

Northwestern British Columbia, Canada

 

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study
and Preliminary Economic Assessment,
NI 43-101 Technical Report

 

Effective Date: August 8, 2022

 

 

Prepared by:

 

Hassan Ghaffari, P.Eng., Tetra Tech, Inc.  
Henry Kim, P. Geo., Wood Canada Limited
Jianhui (John) Huang, Ph.D., P.Eng., Tetra Tech, Inc.
James H. Gray, P.Eng., Moose Mountain Technical Services
Derek Kinakin, M.Sc., P.Geo., P.G., BGC Engineering Inc.
David Willms, P.Eng., Klohn Crippen Berger Ltd.
Neil Brazier, P.Eng., WN Brazier Associates Inc.
Rolf Schmidt, P. Geo., ERM Consultants Canada Ltd.
Ross Hammett, Ph.D., P.Eng., WSP Golder Inc.

 

 

  

 

Important Notice

 

 

This report was prepared as National Instrument 43-101 Technical Report for Seabridge Gold Inc. (Seabridge) by Tetra Tech, Inc., Moose Mountain Technical Services, WSP Golder Inc., BGC Engineering Inc., Klohn Crippen Berger Ltd, ERM Consultants Canada Ltd, WN Brazier Associates Inc., and Wood Canada Limited, collectively the Report Authors. The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in the Report Authors’ services, based on i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by Seabridge subject to the respective terms and conditions of its contracts with the individual Report Authors. Those contracts permit Seabridge to file this report as a Technical Report with Canadian Securities Regulatory Authorities pursuant to provincial and territorial securities law. Except for the purposes legislated under Canadian provincial and territorial securities law, any other use of this report by any third party is at that party’s sole risk.

 

Seabridge Gold Inc.ii221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

CONTENTS

 

 

1.0 Summary 1-1
  1.1 Introduction 1-1
  1.2 Property Location and Ownership 1-2
  1.3 Accessibility, Climate, Local Resources, Physiography, and Infrastructure 1-3
  1.4 History 1-3
  1.5 Geological Setting and Mineralization 1-4
  1.6 Mineral Resources 1-5
  1.7 Mining Methods 1-7
    1.7.1 Mineral Reserve Estimate 1-7
    1.7.2 Mine Production Plan 1-9
  1.8 Mineral Processing and Metallurgical Testing 1-9
  1.9 2022 PFS Infrastructure 1-11
    1.9.1 Geohazards 1-11
    1.9.2 Tailings Management 1-11
    1.9.3 Mine Site Water Management 1-12
    1.9.4 Tunneling 1-12
    1.9.5 Mine To Mill Ore Transport System 1-13
  1.10 Environmental Studies, Permitting, and Social or Community Impact 1-13
    1.10.1 Benefit Agreements Tahltan Nation 1-13
    1.10.2 Closure and Reclamation 1-13
  1.11 2022 PFS Capital Cost Estimate 1-13
  1.12 2022 PFS Operating Cost Estimate 1-15
  1.13 2022 PFS Economic Evaluation 1-15
    1.13.1 Sensitivity Analysis 1-17
  1.14 2022 Preliminary Economic Assessment 1-18
    1.14.1 Introduction 1-18
    1.14.2 Mining Methods 1-18
    1.14.3 Open Pit Mining 1-18
    1.14.4 Underground Mining 1-18
    1.14.5 Mine Production Plan 1-19
    1.14.6 Recovery Methods 1-20
    1.14.7 2022 PEA Capital and Operating Costs 1-20
    1.14.8 Economic Analysis 1-21
  1.15 Conclusions 1-22
  1.16 Recommendations 1-22
2.0 Introduction 2-1
  2.1 Overview 2-1
  2.2 Terms of Reference 2-1
    2.2.1 Mineral Resource Estimate Update 2-1
    2.2.2 2022 PFS and 2022 PEA Press Releases 2-2

 

Seabridge Gold Inc.iii221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

  2.3 Sources of Information 2-2
  2.4 Effective Dates 2-2
  2.5 Qualified Persons 2-2
  2.6 Personal Inspections 2-2
3.0 Reliance on Other Experts 3-1
4.0 Property Description and Location 4-1
  4.1 Mineral Tenure 4-1
  4.2 Royalties 4-9
    4.2.1 KSM Property Royalties 4-9
    4.2.2 Sprott and Teachers Royalties 4-9
5.0 Accessibility, Climate, Infrastructure, Local Resources and Physiography 5-1
6.0 History 6-1
  6.1 Exploration History 6-1
  6.2 History of Production 6-6
7.0 Geological Setting and Mineralization 7-1
  7.1 Geological Setting 7-1
  7.2 Mineralization 7-3
    7.2.1 Kerr Zone 7-3
    7.2.2 Sulphurets Zone 7-7
    7.2.3 Mitchell Zone 7-13
    7.2.4 East Mitchell Zone 7-18
    7.2.5 Iron Cap Zone 7-21
8.0 Deposit Types 8-1
9.0 Exploration 9-1
  9.1 2011 Geophysical Exploration Program 9-1
    9.1.1 Results of 2011 Geophysical Program 9-1
  9.2 2013 Geophysical Exploration Program 9-1
    9.2.1 Results of 2013 Geophysical Program 9-2
  9.3 2014 Geophysical Exploration Program 9-3
    9.3.1 Results of 2014 Geophysical Programs 9-3
  9.4 2015 Geophysical Exploration Program 9-3
  9.5 2019 Geophysical Program 9-4
  9.6 East Mitchell Exploration 9-5
10.0 Drilling 10-1
  10.1 Introduction 10-1
  10.2 Type and Extent of drilling 10-4
  10.3 Drilling Procedures 10-10
  10.4 Recent Drilling – Mitchell and East Mitchell 10-11
  10.5 QP Comments Regarding Drilling and Sampling Factors 10-12

 

Seabridge Gold Inc.iv221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

11.0 Sample Preparation, Analysis and Security 11-1
  11.1 Introduction 11-1
  11.2 KSM Sample Preparation Methods and Procedures 11-1
    11.2.1 Statement on Sample Preparation Personnel 11-1
    11.2.2 Sample Preparation and Dispatch 11-1
    11.2.3 Pre-2012 Analytical Procedures 11-2
    11.2.4 Post-2012 Analytical Procedures 11-3
  11.3 Summary of the Nature, Extent, and Results of Quality Control Procedures 11-6
    11.3.1 Pre-2012 Quality Control Procedures 11-9
    11.3.2 Post-2012 Quality Control Procedures 11-10
  11.4 QP’s Opinion 11-10
12.0 Data Verification 12-1
  12.1 Introduction 12-1
  12.2 Data Verification by Wood Qualified Persons 12-1
  12.3 QP’s Opinion 12-2
    12.3.1 Drill Hole Data Verification 12-2
13.0 Mineral Processing and Metallurgical Testing 13-1
  13.1 Introduction 13-1
  13.2 Summary of Metallurgical Test Programs 13-4
  13.3 Summary of Initial Test Work 1989–1991 13-5
  13.4 Summary of Test Work 2007–2022 13-6
    13.4.1 Test Programs 13-6
    13.4.2 Baseline Test Process Flowsheet and Conditions 13-6
    13.4.3 Mitchell Deposit Test Results 13-7
    13.4.4 East Mitchell Deposit Test Results 13-17
    13.4.5 Sulphurets Deposit Test Results 13-24
    13.4.6 Upper Kerr Zone Metallurgical Test Results 13-26
    13.4.7 Lower Kerr Zone Metallurgical Test Results 13-28
    13.4.8 Iron Cap Deposit Metallurgical Test Results 13-34
    13.4.9 Flotation Concentrate Assay (2007–2022) 13-42
    13.4.10 Ancillary Tests 13-44
  13.5 Conclusions 13-49
  13.6 Metallurgical Performance Projection 13-50
    13.6.1 Mitchell, East Mitchell, Sulphurets, Upper Kerr, and Upper Iron Cap 13-51
    13.6.2 Metallurgical Performance Projection – Lower Kerr and Lower Iron Cap 13-55
14.0 Mineral Resource Estimates 14-1
  14.1 Kerr Deposit 14-1
    14.1.1 Grade Distribution – Kerr Deposit 14-1
    14.1.2 Assay Grade Capping – Kerr Deposit 14-1
    14.1.3 Drill Hole Compositing – Kerr Deposit 14-2
    14.1.4 Geological Constraints - Kerr Deposit 14-2
    14.1.5 Variography – Kerr Deposit 14-2

 

Seabridge Gold Inc.v221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

    14.1.6 Grade Estimation Parameters – Kerr Deposit 14-3
    14.1.7 Grade Model Verification - Kerr Deposit 14-3
    14.1.8 Resource Classification – Kerr Deposit 14-9
  14.2 Sulphurets Deposit 14-9
    14.2.1 Grade Distribution – Sulphurets Deposit 14-9
    14.2.2 Assay Grade Capping – Sulphurets Deposit 14-9
    14.2.3 Drill Hole Compositing – Sulphurets Deposit 14-10
    14.2.4 Geological Constraints – Sulphurets Deposit 14-10
    14.2.5 Variography – Sulphurets Deposit 14-10
    14.2.6 Grade Estimation Parameters – Sulphurets Deposit 14-10
    14.2.7 Model Validation – Sulphurets Deposit 14-11
    14.2.8 Resource Classification – Sulphurets Deposit 14-17
  14.3 Mitchell and East Mitchell Deposit 14-17
    14.3.1 Metal Distribution – Mitchell and East Mitchell Deposit 14-17
    14.3.2 Assay Grade Capping – Mitchell Deposit 14-18
    14.3.3 Drill Hole Compositing – Mitchell Deposit 14-19
    14.3.4 Variography – Mitchell and East Mitchell Deposit 14-19
    14.3.5 Grade Estimation Parameters – Mitchell Deposit 14-19
    14.3.6 Grade Model Validation – Mitchell Deposit 14-22
    14.3.7 Density Estimation 14-33
    14.3.8 Resource Classification – Mitchell Deposit 14-34
  14.4 Iron Cap Deposit 14-35
    14.4.1 Grade Distribution – Iron Cap Deposit 14-35
    14.4.2 Assay Grade Capping – Iron Cap Deposit 14-36
    14.4.3 Drill Hole Compositing – Iron Cap Deposit 14-36
    14.4.4 Geological Constraints – Iron Cap Deposit 14-36
    14.4.5 Variography – Iron Cap Deposit 14-36
    14.4.6 Grade Estimation Parameters – Iron Cap Deposit 14-36
    14.4.7 Grade Model Verification – Iron Cap Deposit 14-37
    14.4.8 Resource Classification – Iron Cap Deposit 14-43
  14.5 Bulk Density 14-43
  14.6 Resource Criteria 14-43
  14.7 Summary of KSM Mineral Resources 14-44
  14.8 General Discussion 14-46
15.0 Mineral Reserve Estimates 15-1
  15.1 Introduction 15-1
  15.2 Open Pit Reserve Parameters 15-1
  15.3 Mineral Reserves 15-1
  15.4 Factors that could affect the mineral reserve estimate 15-2
  15.5 Comments on Section 15.0 15-3
16.0 Mining Methods 16-1
  16.1 Introduction 16-1
  16.2 Open Pit Mining Operations 16-1
    16.2.1 Introduction 16-1
    16.2.2 Mining Datum 16-1

 

Seabridge Gold Inc.vi221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

    16.2.3 Open Pit Mine Planning 3D Block Model 16-1
    16.2.4 Pit Slope Design Angles 16-3
    16.2.5 Economic Pit Limits, Pit Designs 16-7
    16.2.6 Detailed Pit Designs 16-12
    16.2.7 Open Pit Mine Plan 16-17
    16.2.8 Open Pit Production 16-22
    16.2.9 Open Pit Mine Operations 16-25
    16.2.10 Mine Closure and Reclamation 16-27
    16.2.11 Open Pit Mine Equipment Parameters 16-27
  16.3 Mine Production Schedule 16-31
17.0 Recovery Methods 17-1
  17.1 Introduction 17-1
  17.2 Major Process Design Criteria 17-3
  17.3 Process Plant Description 17-5
    17.3.1 Primary Crushing 17-5
    17.3.2 Coarse Ore Transport From Mitchell Site to Treaty Site 17-5
    17.3.3 Coarse Material Handling 17-6
    17.3.4 Secondary Crushing 17-6
    17.3.5 Tertiary Crushing Material Conveyance/Storage 17-7
    17.3.6 Primary Grinding 17-7
    17.3.7 Copper, Gold and Molybdenum Flotation 17-8
    17.3.8 Concentrate Dewatering 17-9
    17.3.9 Gold Recovery From Gold-bearing Pyrite Products 17-10
    17.3.10 Treatment of Leach Residue 17-13
    17.3.11 Tailing Management 17-14
    17.3.12 Reagents Handling 17-14
    17.3.13 Water Supply 17-15
    17.3.14 Air Supply 17-16
    17.3.15 Assay and Metallurgical Laboratory 17-17
    17.3.16 Process Control and Instrumentation 17-17
  17.4 Yearly Production Projection 17-18
18.0 Project Infrastructure 18-1
  18.1 Site Layout 18-1
  18.2 Tailings, Mine Rock, and Water Management 18-4
    18.2.1 Introduction 18-4
    18.2.2 Mine Site Characterization 18-5
    18.2.3 TMF and PTMA Site Characterization 18-8
    18.2.4 Rock Storage Facilities 18-11
    18.2.5 Mine Site Water Management 18-11
    18.2.6 Water Treatment 18-16
    18.2.7 Tailings Management Facility Design 18-18
  18.3 Tunnels 18-28
    18.3.1 Mitchell-Treaty Tunnels 18-30
  18.4 Mine To Mill Ore Transport System 18-33
    18.4.1 MTT Freight and Personnel Transport 18-37

 

Seabridge Gold Inc.vii221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

  18.5 Site Roads 18-39
    18.5.1 Road Width 18-39
  18.6 Ancillary Buildings 18-40
    18.6.1 Treaty OPC 18-41
    18.6.2 Mine Site 18-42
  18.7 Sewage 18-42
  18.8 Communications System 18-43
  18.9 Fresh and Potable Water Supply 18-43
  18.10 Power Supply and Primary Distribution 18-44
    18.10.1 Northwest Transmission Line 18-44
    18.10.2 Treaty Creek Switching Station 18-45
    18.10.3 Transmission Line Extension To KSM 18-46
    18.10.4 System Studies 18-47
    18.10.5 Electric Utility Requirements, Tariffs, and Cost of Electric Power 18-48
    18.10.6 Treaty Plant Main Substation (FLT1) 18-49
    18.10.7 138 kV Cable 18-50
    18.10.8 Mitchell Substation FLT2 18-50
    18.10.9 Site Power Distribution 18-51
    18.10.10 Mine Power 18-51
    18.10.11 Trolley assist 18-51
    18.10.12 Construction and Standby Power 18-52
    18.10.13 Energy Recovery and Self Generation 18-53
  18.11 Treaty OPC and Mine Site Secondary Electrical Power Distribution and Utilization 18-54
    18.11.1 Mine and Plant Power Consumption 18-54
    18.11.2 Power Distribution – Treaty Plant Main Substation FLT1 18-54
    18.11.3 Mitchell Substation FLT2 18-55
  18.12 Permanent and Construction Access Roads 18-56
    18.12.1 Route Descriptions 18-58
    18.12.2 Road Design Requirements 18-60
    18.12.3 Design Progress 18-61
  18.13 Logistics 18-63
  18.14 Preliminary Construction Execution Plan 18-64
    18.14.1 Introduction 18-64
    18.14.2 Early Works Plan 18-65
    18.14.3 Construction Scope 18-66
    18.14.4 Construction Schedule 18-68
    18.14.5 Engineering and Procurement 18-69
    18.14.6 Construction Management 18-70
    18.14.7 Construction Supervision and Contractor Management 18-70
    18.14.8 Contracting Packaging and Strategy Overview 18-70
    18.14.9 Site Organization Structure 18-70
    18.14.10 Environmental and Community Affairs 18-71
    18.14.11 Pre-commissioning/Commissioning 18-72
  18.15 Owner’s Implementation Plan 18-72

 

Seabridge Gold Inc.viii221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

19.0 Market Studies and Contracts 19-1
  19.1 Copper Concentrate 19-1
    19.1.1 Marketability 19-1
    19.1.2 Smelting Terms 19-1
  19.2 Molybdenite Concentrate 19-5
    19.2.1 Smelting Charge 19-5
  19.3 Gold and Silver Doré 19-6
20.0 Environmental Studies, Permitting, and Social or Community Impact 20-1
  20.1 Licensing and Permitting 20-1
    20.1.1 Provincial EA Process 20-2
    20.1.2 Federal Process 20-3
    20.1.3 Provincial Permits 20-4
    20.1.4 Federal Permits 20-5
    20.1.5 benefits agreements 20-6
    20.1.6 Biophysical Setting 20-7
    20.1.7 Economic, Social, and Cultural Setting 20-9
  20.2 Water Management 20-11
    20.2.1 Overview of Water Management 20-11
    20.2.2 Summary of Water Management Plan 20-13
  20.3 Waste Management 20-14
    20.3.1 Tailing Management Facility Management and Monitoring Plan 20-14
    20.3.2 Best Available Tailings Technology Assessment 20-15
    20.3.3 Waste Rock Management 20-15
    20.3.4 Domestic and Industrial non-hazardous and hazardous Waste Management 20-16
  20.4 Air Quality Management including Greenhouse Gases 20-17
  20.5 Environmental Management System 20-17
  20.6 Closure and Reclamation 20-18
    20.6.1 Closure and Reclamation Objectives 20-18
    20.6.2 Soil Handling Plan 20-19
    20.6.3 Closure and Reclamation Planning 20-19
21.0 Capital and Operating Cost Estimates 21-1
  21.1 Initial Capital Costs 21-1
    21.1.1 Exclusions 21-2
    21.1.2 Direct Costs 21-3
    21.1.3 Indirect Costs 21-5
    21.1.4 Owner’s Costs 21-6
    21.1.5 Contingency 21-6
  21.2 Sustaining Capital Costs 21-6
    21.2.1 Mine Site 21-7
    21.2.2 Open Pit Mining 21-7
    21.2.3 Mine Site Water Treatment 21-8
    21.2.4 Process 21-8
    21.2.5 Tailing Management Facility 21-8

 

Seabridge Gold Inc.ix221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

    21.2.6 Other Sustaining Capital Costs 21-9
  21.3 Operating Costs 21-9
    21.3.1 Open Pit Mine Operating Costs 21-11
    21.3.2 Process Operating Costs 21-12
    21.3.3 TMF Dam Management Operating Costs 21-16
    21.3.4 Mine Site Water Management Costs 21-16
    21.3.5 General and Administrative 21-16
    21.3.6 Site Services 21-17
22.0 Economic Analysis 22-1
  22.1 Introduction 22-1
  22.2 Forward-looking Statements 22-2
  22.3 Pre-Tax Model 22-3
    22.3.1 Financial Evaluations: NPV and IRR 22-3
    22.3.2 Metal Price Scenarios 22-4
  22.4 Post-tax Financial Evaluations 22-5
    22.4.1 Canadian Federal and BC Provincial Income Tax Regime 22-5
    22.4.2 BC Mineral Tax Regime 22-5
    22.4.3 Taxes and Post-tax Financial Results 22-6
  22.5 Sensitivity Analysis 22-10
  22.6 Royalties 22-11
  22.7 Smelter Terms 22-11
  22.8 Miscellaneous Costs and Charges 22-12
23.0 Adjacent Properties 23-1
24.0 Other Relevant Data and Information 24-1
  24.16 Mining Methods 24-3
    24.16.1 Net Smelter Return Block Model 24-3
    24.16.2 Open Pit Mining Method 24-3
    24.16.3 Iron Cap Mining Methods 24-5
    24.16.4 Kerr Mining Methods 24-17
    24.16.5 Mine Production Schedule (Open Pit and Underground Combined) 24-31
  24.17 Recovery Methods 24-33
    24.17.1 Process Plant 24-34
    24.17.2 flowsheet description 24-34
  24.18 Project Infrastructure 24-38
    24.18.1 On-Site Infrastructure 24-38
    24.18.2 Off-site Infrastructure 24-40
    24.18.3 Iron Cap Project Infrastructure 24-42
    24.18.4 Kerr Project Infrastructure 24-44
  24.19 Market Studies and Contracts 24-47
  24.20 Environmental Studies, Permitting and Social or Community Impact 24-50
  24.21 Capital and Operating Cost Estimates 24-52
    24.21.1 Capital Cost Estimate 24-52
    24.21.2 Operating Cost Estimate 24-54
  24.22 Economic Analysis 24-58
    24.22.1 Methodology Used 24-59
    24.22.2 Financial Model Parameters 24-60

 

Seabridge Gold Inc.x221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

25.0 Interpretations and Conclusions 25-1
  25.1 Introduction 25-1
  25.2 2022 Prefeasibility Study Conclusions 25-1
    25.2.1 Exploration and Mineral Resources 25-1
    25.2.2 2022 PFS Mineral Reserves 25-2
    25.2.3 Mining Methods 25-2
    25.2.4 Recovery Methods 25-3
    25.2.5 2022 PFS Project Infrastructure 25-3
    25.2.6 2022 PFS Economic Analysis 25-5
  25.3 2022 Preliminary Economic Assessment Conclusions 25-7
    25.3.1 Mining Methods 25-7
    25.3.2 Recovery Methods 25-8
    25.3.3 2022 PEA Project Infrastructure 25-8
    25.3.4 2022 PEA Economic Analysis 25-9
  25.4 2022 Prefeasibility Study Risks 25-10
  25.5 2022 Preliminary Economic Assessment Risks 25-15
  25.6 Mineral Resource Risks 25-17
26.0 Recommendations 26-1
  26.1 Introduction 26-1
  26.2 Data Collection to Complete A Feasibility Study 26-1
    26.2.1 Pit Slope Geotech 26-1
    26.2.2 Rock Storage Facilities Geotech 26-2
    26.2.3 Metallurgical Testing 26-3
    26.2.4 Water Management 26-4
    26.2.5 TMF Area 26-4
    26.2.6 Tunnels 26-5
    26.2.7 Site Infrastructure Geotech 26-5
27.0 References   27-1

 

Seabridge Gold Inc.xi221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

  

List of Tables

 

 

Table 1.1

KSM Mineral Resources 1-6
Table 1.2 KSM Proven and Probable Mineral Reserves 1-8
Table 1.3 2022 PFS Initial Capital Cost Summary 1-14
Table 1.4 2022 PFS LOM Average Unit Operating Costs 1-15
Table 1.5 2022 PFS Summary of the Pre-tax Economic Evaluations 1-16
Table 1.6 2022 PFS Summary of the Post-tax Economic Evaluations 1-16
Table 1.7 Production Tonnes and Grade in 2022 PEA Mine Plan 1-19
Table 1.8 2022 PEA Capital Cost Estimate Summary 1-20
Table 1.9 2022 PEA LOM Average Unit Operating Costs 1-21
Table 1.10 2022 PEA Economic Results 1-22
Table 4.1 KSM Mineral Claims and Leases 4-5
Table 4.2 KSM Placer Claims 4-6
Table 4.3 Seabee/Tina KSM Claims 4-7
Table 10.1 KSM Core Drilling Through 2021 10-2
Table 10.2 Drilling by Company Through 2021 10-2
Table 10.3 KSM Drill Hole Summary by Area and Company Through 2021 10-3
Table 10.4 2021 Drill Hole Assay Results 10-12
Table 11.1 ICP Detection Limits – Pre-2012 Data 11-3
Table 11.2 ICP Detection Limits – Post 2012 11-4
Table 11.3 Aqua regia digestion, ICP-MS and ICP-AES analysis – 2019 11-5
Table 11.4 4-acid detection limits by ICP-AES – 2021 11-6
Table 11.5 Summary of KSM Control Samples Submitted Thru Time 11-6
Table 11.6 ¼ Core Duplicate Sample Statistics (2006-2021) 11-7
Table 13.1 Typical Mineralogical Characteristics and Average Copper-to-Gold Grade Ratios in Potential Mill Feed 13-1
Table 13.2 Average Ball Mill Grindability and Abrasion Index 13-2
Table 13.3 Locked Cycle Flotation Test Result Summary 13-3
Table 13.4 Metallurgical Test Work Programs 13-4
Table 13.5 Test Samples – Mitchell (2007–2022) 13-8
Table 13.6 JK SimMet Simulation Results (60,000 t/d SABC Circuit, 2008) 13-8
Table 13.7 Locked Cycle Test Results – Mitchell 13-12
Table 13.8 Locked Cycle Test Results – Blended Samples (Mitchell and Other Deposits) 13-13
Table 13.9 Cu-Mo Separation LCT Results, 2010 13-14
Table 13.10 Cyanidation Test Results on LCT Products – Mitchell 13-15
Table 13.11 Test Samples – East Mitchell (2011 and 2021) 13-18
Table 13.12 LCT Results – East Mitchell 13-21
Table 13.13 Direct Cyanidation _ East Mitchell Upper Zone Master Composite 13-22
Table 13.14 Cyanidation Test Results on LCT Products – East Mitchell 13-23
Table 13.15 Test Samples – Sulphurets (2009-2012) 13-24
Table 13.16 LCT Results – Sulphurets 13-25
Table 13.17 Cyanidation Test Results – Flotation LCT Products, Sulphurets, 2009–2011 13-25
Table 13.18 Metal Contents of Composites – Upper Kerr, 2010 (G&T) 13-26
Table 13.19 LCT Results – Upper Kerr (G&T) 13-27
Table 13.20 Cyanidation Test Results on LCT Products – Upper Kerr (G&T) 13-28
Table 13.21 Test Samples – Lower Kerr (2012–2017) 13-28
Table 13.22 Bond Ball Mill Work Index Test Results – Lower Kerr, 2017 (ALS KM5063) 13-29
Table 13.23 Mineral Composition Data – Lower Kerr 2016/2017 (ALS) 13-29
Table 13.24 Flotation LCT Results – Lower Kerr 13-31

 

Seabridge Gold Inc.xii221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

Table 13.25 Preliminary Cyanidation Test Results – Lower Kerr (before 2016) 13-32
Table 13.26 Preliminary Cyanidation Test Results – Lower Kerr 13-32
Table 13.27 Test Samples – Iron Cap 13-34
Table 13.28 Bond Ball Mill Work Index Test Results – Iron Cap (ALS) 13-35
Table 13.29 Mineral Composition Data – Iron Cap 2017-2020 (ALS) 13-36
Table 13.30 Flotation Locked Cycle Test Results – Iron Cap 2010 -2020 13-38
Table 13.31 Cyanidation Test Results on LCT Products – Iron Cap 13-40
Table 13.32 Preliminary Cyanidation Test Results – Iron Cap (ALS 2017 - 2020) 13-41
Table 13.33 Flotation Concentrate Assay from Different Deposit Samples 13-43
Table 13.34 Flotation Concentrate Assay from Blended Head Samples 13-44
Table 13.35 Chemical Analysis of Cyanide Recovery Test Solution and Cyanide Destruction Pulp 13-45
Table 13.36 Cyanide Recovery Test Results – AVR 13-45
Table 13.37 Cyanide Destruction Test Results – 2009/2010 (SGS) 13-47
Table 13.38 Recommended Conventional Thickener Operating Parameters – 2009 (Pocock) 13-48
Table 13.39 Recommended High Rate Thickener Operating Parameters – 2009 (Pocock) 13-48
Table 13.40 Filtration Test Results – 2009 (Pocock) 13-48
Table 13.41 Cu-Au Flotation Concentrate Grade Versus Cu Head Grade 13-51
Table 13.42 Cu-Au Flotation Concentrate – Metal Recovery Projections 13-51
Table 13.43 Au-Ag Doré – Cyanide Leach Metal Recovery Projections* 13-53
Table 13.44 Mo Flotation Concentrate Metal Recovery and Grade (Molybdenum Concentrate Grade = 50%) 13-55
Table 13.45 Cu-Au Flotation Concentrate –Metal Recovery Projections 13-56
Table 13.46 Au-Ag Doré – Cyanide Leach Metal Recovery Projections 13-57
Table 13.47 Mo Flotation Concentrate Metal Recovery and Grade (Molybdenum Concentrate Grade = 50%) 13-57
Table 14.1 Distribution of Gold by Lithology and Estimation Domain – Mitchell and East Mitchell Deposit 14-17
Table 14.2 Distribution of Copper by Lithology and Estimation Domain – Mitchell and East Mitchell Deposit 14-18
Table 14.3 Capped Silver Assay Intervals and Grades 14-18
Table 14.4 Estimation Parameters for Gold, Copper, Silver and Molybdenum 14-20
Table 14.5 Validation of Global Bias and Metal Reduction 14-27
Table 14.6 Change of Support for Gold at Mitchell and East Mitchell (Main Zone) 14-31
Table 14.7 Grade-Tonnage Metrics for Mitchell Gold Grade Estimates and Theoretical HERCO Distribution 14-32
Table 14.8 Grade-Tonnage Metrics for East Mitchell Gold Grade Estimates and Theoretical HERCO Distribution 14-33
Table 14.9 Drillhole Spacing Study Results for Mitchell and East Mitchell 14-35
Table 14.10 Key Mineral Resource Parameters for Kerr, Sulphurets and Iron Cap 14-43
Table 14.11 Key Mineral Resource Parameters for Mitchell and East Mitchell 14-44
Table 14.12 KSM Mineral Resources 14-45
Table 15.1 Proven and Probable Reserves 15-2
Table 16.1 Major Smelter Terms Used in the NSR Calculation 16-2
Table 16.2 Metal Prices and Resultant NSP 16-3
Table 16.3 LG Pit Limit Primary Assumptions 16-8
Table 16.4 Production Schedule Assumptions 16-19
Table 16.5 Major Equipment Requirements 16-28
Table 16.6 Blasting Assumptions 16-28
Table 16.7 Open Pit Production Drilling Assumptions 16-29
Table 16.8 Mine Support Equipment Fleet 16-30

 

Seabridge Gold Inc.xiii221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

Table 16.9 Summarized Production Schedule 16-32
Table 17.1 Major Design Criteria 17-3
Table 17.2 Projected Metallurgical Performance 17-19
Table 17.3 Projected Copper Concentrate Quality 17-20
Table 18.1 Climate Data for the Mine Site (Sulphurets Creek Climate Station) 18-7
Table 18.2 Climate Data for the TMF (Teigen Creek Climate Station)1 18-10
Table 18.3 Temporary Water Treatment Plant Locations 18-16
Table 18.4 Average Annual Reagent Consumption for the HDS WTP 18-17
Table 18.5 Tailings Dam Summary 18-22
Table 18.6 KSM Pre-Production Tunnels Summary 18-29
Table 18.7 KSM Operational Phase Tunnels Summary 18-30
Table 18.8 Mini Hydro and Energy Recovery Power Generation 18-53
Table 18.9 Owner’s Key Activities by Year 18-75
Table 19.1 Benchmark Smelting Terms 19-2
Table 21.1 Initial Capital Cost Summary 21-2
Table 21.2 Mine Site Capital Costs 21-4
Table 21.3 Process-Treaty OPC Capital Cost Estimate 21-5
Table 21.4 Indirect Capital Costs 21-6
Table 21.5 Sustaining Capital Costs 21-7
Table 21.6 Mine Site Sustaining Capital Costs 21-7
Table 21.7 Operating Cost Summary 21-10
Table 21.8 Summary of Process Operating Costs by Deposit 21-14
Table 21.9 Operating Costs per Area of Operation by Deposit 21-15
Table 22.1 Metal Production from the KSM Mine 22-3
Table 22.2 Summary of the Pre-tax Economic Evaluations 22-4
Table 22.3 Component of the Various Taxes for all Scenarios 22-7
Table 22.4 Summary of Post-tax Financial Results 22-7
Table 22.5 2022 PFS Annual Cash Flow for Pre-production Period, Years 1 to 7 and LOM1,2 22-8
Table 24.1 Metal Prices for PEA NSR Calculation 24-3
Table 24.2 Mineral Resources Within the PEA Open Pit Mine Plan 24-4
Table 24.3 Iron Cap Footprint Finder Input Parameters 24-6
Table 24.4 Main Drift Types and Dimensions 24-11
Table 24.5 Vertical Development Dimensions 24-12
Table 24.6 Iron Cap Production Tonnes and Grade in 2022 PEA Mine Plan 24-14
Table 24.7 Peak Mobile Equipment Requirements for Iron Cap 24-16
Table 24.8 Kerr Footprint Finder Input Parameters 24-18
Table 24.9 Main Drift Types and Dimensions 24-25
Table 24.10 Vertical Development Dimensions 24-26
Table 24.11 Kerr Production Tonnes and Grade in 2022 PEA Mine Plan 24-29
Table 24.12 Peak Mobile Equipment Requirements for Kerr 24-30
Table 24.13 Summarized Production Schedule – Open Pit and Underground 24-31
Table 24.14 Capital Cost Estimate Responsibilities by Firm 24-52
Table 24.15 2022 PEA Capital Cost Estimate Summary 24-53
Table 24.16 Operating Cost Estimate Responsibilities by Firm 24-54
Table 24.17 2022 PEA Average Operating Costs 24-55
Table 24.18 Open Pit Mine Operating Costs 24-56
Table 24.19 Underground Mine Operating Costs 24-56
Table 24.20 Process and Tailing Management Operating Costs 24-57
Table 24.21 Water Management Costs 24-58
Table 24.22 2022 PEA Financial Analysis Summary 24-63
Table 25.1 Summary of Major 2022 PFS Pre- and Post-Tax Results by Metal Price Scenario 25-6
Table 25.2 2022 PEA Financial Analysis Summary 25-10
Table 25.3 PFS Risks and Mitigation Measures 25-11
Table 25.4 PEA Risks and Mitigation Measures 25-15
Table 25.5 Mineral Resource Risks 25-17
Table 26.1 Summary of Recommendations and Associated Costs 26-1

 

Seabridge Gold Inc.xiv221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

List of Figures

 

 

Figure 1.1

General Location Map 1-2
Figure 1.2 2022 PFS Mill Feed Production Schedule 1-9
Figure 1.3 2022 PFS Sensitivity Analysis of Post-tax NPV at a 5% Discount Rate 1-17
Figure 1.4 2022 PFS Sensitivity Analysis of Post-tax IRR 1-17
Figure 1.6 2022 PEA Mill Feed Production Schedule 1-20
Figure 4.1 KSM Mineral Claim Map 4-3
Figure 4.2 KSM Placer Claim Map 4-4
Figure 7.1 Geology of the KSM District (Seabridge, 2022) 7-2
Figure 7.2 Geological Map of the Kerr Deposit (Seabridge, 2019). Thin dykes are not displayed. 7-5
Figure 7.3 Four panels of the same vertical E-W section through the Kerr deposit, at 6,258,650N, showing different properties of the deposit. (Seabridge, 2019) 7-6
Figure 7.4 Four panels of the same vertical E-W section through the Kerr deposit, at 6,259,650N, showing different properties of the deposit. (Seabridge, 2019) 7-7
Figure 7.5 Map of the Sulphurets deposit (Seabridge, 2019) 7-10
Figure 7.6 Vertical cross-section through the Sulphurets deposit, looking ENE (Seabridge, 2019) 7-11
Figure 7.7 Vertical cross-section through the Sulphurets deposit, looking NNE (Seabridge, 2019) 7-12
Figure 7.8 Geology Map of the Mitchell and East Mitchell Deposits (Seabridge, 2022) 7-15
Figure 7.9 Vertical section through the Mitchell and East Mitchell Zones (Seabridge, 2022) 7-16
Figure 7.10 Plan view of the Mitchell deposit at 950 masl (Seabridge, 2022) 7-17
Figure 7.11 Vertical section through the East Mitchell Zone (Seabridge, 2022) 7-20
Figure 7.12 Iron Cap geology map. The traces of the Sulphurets Thrust Fault (STF), Johnstone Fault (JF), and Iron Cap Fault (ICF) are shown. (Seabridge, 2019) 7-23
Figure 7.13 Vertical cross-section through the Iron Cap deposit, looking NNE (Seabridge, 2019) 7-24
Figure 7.14 Plan view through the Iron Cap deposit at 1200 m elevation (Seabridge, 2019) 7-25
Figure 10.1 KSM Drill Hole Locations 10-5
Figure 10.2 Drill Hole Locations – Kerr Deposit 10-6
Figure 10.3 Drill Hole Locations – Sulphurets Deposit 10-7
Figure 10.4 Drill Hole Locations – Mitchell Deposit 10-8
Figure 10.5 Drill Hole Locations – Iron Cap Deposit 10-9
Figure 11.1 Duplicate Sample Box Plots 11-8
Figure 11.2 Gold Duplicate Sample Graphs 11-8
Figure 11.3 Copper Duplicate Sample Graphs 11-9
Figure 13.1 Copper and Gold Open Cycle Flotation Variability Test Results (KM2153) 13-10
Figure 13.2 Copper Open Cycle Flotation Performance vs Copper Head Grade at a Concentrate Grade of 25% Copper (KM2153) 13-10

 

Seabridge Gold Inc.xv221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

Figure 13.3 Flotation Performance – Open Circuit Flotation Tests, Mitchell (KM2153) 13-11
Figure 13.4 Copper Recovery vs. Copper Feed – Open Cleaner Circuit Tests (KM2153) 13-11
Figure 13.5 Copper Sulphides and Pyrite Liberation 13-19
Figure 13.6 KM5063 Copper Sulphides Liberation – Mitchell and Lower Kerr 2017 (ALS) 13-30
Figure 13.7 KM5266 Copper Sulphides Liberation – Lower Kerr 2017 (ALS) 13-30
Figure 14.1 Kerr Cross Section 6,259,650 N. – Gold 14-4
Figure 14.2 Kerr Cross Section 6,259,650 N. – Copper 14-5
Figure 14.3 Kerr 850 m Level Plan – Gold 14-6
Figure 14.4 Kerr 850 m Level Plan - Copper 14-7
Figure 14.5 Kerr Gold-Copper Swath Plots by Elevation 14-8
Figure 14.6 Sulphurets Cross Section 20 - Gold 14-12
Figure 14.7 Sulphurets Cross Section 20 – Copper 14-13
Figure 14.8 Sulphurets 1135m Level Plan - Gold 14-14
Figure 14.9 Sulphurets 1135 m Level Plan – Copper 14-15
Figure 14.10 Sulphurets Gold-Copper Swath Plots by Elevation 14-16
Figure 14.11 Mitchell Cross Section of Gold (upper) and Copper (lower) Assay Composite and Block Grade Estimates 14-23
Figure 14.12 East Mitchell Cross Section of Gold (upper) and Copper (lower) Assay Composite and Block Grade Estimates 14-24
Figure 14.13 Mitchell 750 m El Bench Plan of Gold (upper) and Copper (lower) Assay Composite and Block Grade Estimates 14-25
Figure 14.14 950 m El and 1550 m El Bench Plan of Gold (upper) and Copper (lower) Assay Composite and Block Grade Estimates 14-26
Figure 14.15 Swath Plots of Gold, Copper, Silver, and Molybdenum grades for M&I Blocks at Mitchell 14-29
Figure 14.16 Swath Plots of Gold, Copper, Silver, and Molybdenum grades for M&I Blocks at East Mitchell 14-30
Figure 14.17 Grade Tonnage Curve for Mitchel Gold Estimates and Theoretical HERCO Curve 14-31
Figure 14.18 Grade Tonnage Curve for East Mitchell Gold Estimates and Theoretical HERCO Curve 14-32
Figure 14.19 Boxplot of Density Measurements by Lithology 14-33
Figure 14.20 Iron Cap Cross Section 12 – Gold 14-38
Figure 14.21 Iron Cap Cross Section 12 – Copper 14-39
Figure 14.22 Iron Cap 1200 m Level Plan – Gold 14-40
Figure 14.23 Iron Cap 1,200 m Level Plan – Copper 14-41
Figure 14.24 Iron Cap Gold-Copper Swath Plots by Elevation 14-42
Figure 16.1 Mitchell Sensitivity of Ore Tonnes to Pit Size 16-9
Figure 16.2 East Mitchell Sensitivity of Ore Tonnes to Pit Size 16-9
Figure 16.3 Plan View of the KSM LG Pit Limits 16-10
Figure 16.4 Mitchell Economic Pit Limit – North-South Section at East 422950, Viewed from the East 16-11
Figure 16.5 East Mitchell Economic Pit Limit – North-South Section at East 424725, Viewed from the East 16-12
Figure 16.6 Double Lane Haul Road with Trolley Lane 16-13
Figure 16.7 Plan View of Mitchell Pit Phases 16-15
Figure 16.8 Plan View of East Mitchell Pit Phases 16-16
Figure 16-9 Plan View of Sulphurets Pit 16-17
Figure 16.10 End of Pre-production (Year -1) 16-23
Figure 16.11 End of Open Pit Life of Mine 16-24
Figure 16.12 KSM Mill Feed Production Schedule 16-33
Figure 17.1 Simplified Process Flowsheet 17-2

 

Seabridge Gold Inc.xvi221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

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Figure 17.2 Treaty Process Plant Layout 17-4
Figure 18.1 Mine Site Layout after Initial Construction 18-2
Figure 18.2 Ultimate PTMA Layout 18-3
Figure 18.3 Mine Site Ultimate Water Management Facilities 18-12
Figure 18.4 Water Storage Dam Sections 18-15
Figure 18.5 TMF Staging Plan 18-20
Figure 18.6 North Tailings Dam 18-23
Figure 18.7 Saddle and Splitter Tailings Dams 18-24
Figure 18.8 Southeast Tailings Dam 18-25
Figure 18.9 Schematic TMF Water Cycle 18-27
Figure 18.10 MTT General Location and Alignment 18-31
Figure 18.11 MTT Dual Track Plan View 18-35
Figure 18.12 MMT Cross Section (Typical) 18-36
Figure 18.13 Treaty Personnel, Freight, and Fuel Staging and Marshalling 18-38
Figure 18.14 Mitchell Personnel, Freight, and Fuel Staging and Marshalling 18-39
Figure 18.15 NTL Route Map 18-45
Figure 18.16 Proposed Access Roads Network 18-57
Figure 18.17 Construction Schedule Summary 18-69
Figure 18.18 EPCM Organizational Chart 18-71
Figure 20.1 KSM Mine Site Water Management Schematic 20-12
Figure 21.1 Operating Cost Distribution 21-10
Figure 21.2 LOM Average Unit Operating Cost for Open Pit Mining (US$/t Material Mined) – excludes mine pre-production costs 21-12
Figure 22.1 Post-tax Undiscounted Annual and Cumulative Cash Flow 22-6
Figure 22.2 Sensitivity Analysis of Post-tax NPV at a 5% Discount Rate 22-10
Figure 22.3 Sensitivity Analysis of Post-tax IRR 22-11
Figure 24.1 Mine Site General Arrangement (2022 PEA) 24-2
Figure 24.2 Kerr Pit Design – Plan View 24-4
Figure 24.3 Iron Cap Footprint for Lift at 870 m Elevation 24-7
Figure 24.4 Plan View of Iron Cap Mine with Contour 24-9
Figure 24.5 Typical Level Arrangement for Iron Cap 24-10
Figure 24.6 Drawpoint Configuration for Iron Cap 24-11
Figure 24.7 Ventilation Schematic for Iron Cap Mine (Section Looking West) 24-12
Figure 24.8 Iron Cap Mine Production Plan 24-13
Figure 24.9 Proposed Drawpoint Development Sequence for the Iron Cap Mine 24-15
Figure 24.10 Plan View of Kerr 625 m Lift 1 Footprint. 24-19
Figure 24.11 Plan View of Kerr 130 m Lift 2 Footprint. 24-20
Figure 24.12 Kerr Mine Plan View 24-21
Figure 24.13 Kerr Lift 1 (625L Mine Area) General Arrangement 24-22
Figure 24.14 Kerr Lift 2 (130L Mine Area) General Arrangement 24-23
Figure 24.15 Typical Level Arrangement for Kerr Underground 24-23
Figure 24.16 Drawpoint Configuration for Kerr Underground 24-25
Figure 24.17 Ventilation Schematic 24-26
Figure 24.18 Mine Dewatering General Arrangement 24-27
Figure 24.19 Kerr PEA Mine Production Plan 24-28
Figure 24.20 KSM PEA Mill Feed Production Schedule 24-32
Figure 24.21 Overall Process Block Flow Diagram 24-37
Figure 24.22 Post-Tax NPV 5% Sensitivity Analysis 24-64
Figure 24.23 Post-Tax IRR Sensitivity Analysis 24-64

 

Seabridge Gold Inc.xvii221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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Glossary

 

 

Units of Measure

 

ampere A
annum (year) a
bank cubic metres bcm
billion tonnes Bt
billion B
centimetre cm
circular mils cmils
Coefficients of Variation CVs
cubic centimetre cm3
cubic metre m3
day d
days per week d/wk
days per annum (year) d/a
dead weight tonnes DWT
degree °
degrees Celsius °C
diameter ø
dollar (American) US$
dollar (Canadian) Cdn$
dry metric tonne dmt
foot ft
gallon gal
gallons per minute (US) gpm
gigawatt hours Gwh
gigawatt GW
gram g
grams per litre g/L
grams per tonne g/t
gravitational constant g
greater than
hectare ha
horsepower hp
hour h
hours per day h/d
hours per week h/wk
hours per annum (year) h/a
inch in
kilo (thousand) k
kilogram kg

 

Seabridge Gold Inc.xviii221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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kilograms per cubic metre kg/m3
kilograms per day kg/d
kilograms per hour kg/h
kilograms per square metre kg/m2
kilometre km
kilometres per hour km/h
kilonewton kN
kilopascal kPa
kilotonne kt
kilovolt kV
kilowatt hour kWh
kilowatt hours per tonne kWh/t
kilowatt hours per annum (year) kWh/a
kilowatt kW
less than
litre L
litres per hour L/h
litres per second L/s
Megaannum (million years) Ma
megavolt-ampere MVA
megawatt MW
metre m
metres above sea level masl
metres per second m/s
metres per annum (year) m/a
microns µm
milligram mg
milligrams per litre mg/L
millilitre mL
millimetre mm
million M
million tonnes Mt
minute (plane angle)
minute (time) min
month mo
ounce oz
parts per billion ppb
parts per million ppm
Pascal Pa
percent %
pound(s) lb
pound(s) per square inch psi
revolutions per minute rpm
second (plane angle)
second (time) s

 

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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specific gravity SG
square centimetre cm2
square kilometre km2
square metre m2
square millimetre mm2
three-dimensional 3D
tonne (1,000 kg) (metric ton) t
tonnes per day t/d
tonnes per hour t/h
tonnes per annum (year) t/a
troy ounce oz
volt V
week wk
weight/weight w/w
wet metric tonne wmt

 

Abbreviations and Acronyms

 

AACE® International AACE®
abrasion index Ai
acid rock drainage ARD
acidification, volatilization of hydrogen cyanide gas, and re-neutralization AVR
Application Information Requirements AIR
Allnorth Engineering Consulting AIlnorth
ALS Canada Ltd. ALS
alternating current AC
aluminum oxide Al2O3
ammonium nitrate-fuel oil ANFO
antimony Sb
arsenic As
atomic absorption AA
atomic emission spectroscopy AES
ball mill work index BMWi
Barrick Gold Corp. Barrick
Basal Shear Zone BSZ
BC Environmental Assessment Act BCEAA
BC Mineral Inventory Minfile
best available tailings technology BATT
BGC Engineering Inc. BGC
BioteQ Environmental Technologies Inc. BioteQ
bismuth Bi
BQE Water BQE
Brenda Mines Ltd. Metallurgical Laboratory BMML
British Columbia BC
British Columbia Environmental Assessment Office BCEAO

 

Seabridge Gold Inc.xx221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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British Columbia Utilities Commission BCUC
Bulk Mineral Analysis with Liberation BMAL
cadmium Cd
calcium oxide CaO
Canadian Dam Association CDA
Canadian Development Expense CDE
Canadian Environmental Assessment Act CEAA
Canadian Exploration Expense CEE
Canadian Institute of Mining, Metallurgy and Petroleum CIM
Canadian National Railroad CNR
carbon-in-leach CIL
carboxymethyl cellulose CMC
Caro’s acid H2SO5
CDN Resource Laboratories Ltd. CDN Resource
chalcopyrite Cp
chlorine Cl
closed-circuit television CCTV
coarse ore stockpile COS
Coastech Research Inc. CRI
cobalt Co
common voltage reference CVR
Construction Diversion Tunnel CDT
conventional counter-current decantation CCD
copper Cu
copper sulphate CuSO4
cost breakdown structure CBS
Cost, Insurance and Freight – Free Out CIF-FO
Coulter Creek Access Road CCAR
counter current decantation CCD
cross-linked polyethylene XLPE
Cumulative Tax Credit Account CTCA
Delegation of Authority Guideline DOAG
Demand Side Management DSM
direct current DC
direct cyanide leaching DCN
discounted cash flow DCF
Distributed Control System DCS
east E
EBC Inc. EBC
economic, social, and cultural impact assessment ESCIA
effective grinding length EGL
engineering, procurement, construction management EPCM
environmental assessment EA
environmental impact statement EIS
Environmental Management System EMS

 

Seabridge Gold Inc.xxi221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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ERM Consultants Canada Inc. ERM
Esso Minerals Canada Ltd Esso Minerals
Feasibility Study FS
Fisheries and Oceans Canada DFO
fluorine F
Footprint Finder FF
Free Carrier FCA
front-end loader FEL
G&T Metallurgical Services Ltd. G&T
gangue Gn
gas insulated GIS
general and administrative G&A
general mine expense GME
global positioning system GPS
gold Au
gold equivalent AuEQ
WSP Golder Inc. WSP Golder
Goods and Services Tax GST
Granduc Mines Ltd. Granduc
Granmac Services Ltd. Granmac
gross vehicle weight GVW
harmful alteration, disruption or destruction HADD
Harmonized Sales Tax HST
Hazen Research Inc. Hazen
health, safety and security HS&S
heating, ventilation, and air conditioning HVAC
height of draw HOD
hematite He
high-density polyethylene HDPE
high-density sludge HDS
high-pressure grinding roll HPGR
hydrochloric acid HCl
hydrogen peroxide H2O2
Impact Benefits Agreement IBA
Independent Geotechnical Review Board IGRB
Independent Power Producer IPP
induced polarization IP
inductively coupled plasma ICP
Integrated Remote Operating Centre IROC
internal rate of return IRR
International Electrotechnical Commission IEC
International Organization for Standardization ISO
inverse distance weighting IDW
Iron Cap Fault ICF
iron Fe

 

Seabridge Gold Inc.xxii221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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joint venture JV
Kerr-Sulphurets-Mitchell KSM
KSMCo KSM Mining ULC
Klohn Crippen Berger Ltd. KCB
Land Resource Management Plan LRMP
lead Pb
Lerchs-Grossmann LG
Life of mine LOM
Light Detection and Ranging LiDAR
load factor LF
load-haul-dump LHD
local study area LSA
locked cycle tests LCT
magnesium oxide MgO
magnetite Ma
magneto telluric MT
maintenance and repair contracts MARC
manganese oxide MnO
McElhanney Consulting Services Inc. McElhanney
McTagg Diversion Tunnels MTDT
mercury Hg
metabisulphite MBS
Metal and Diamond Mining Effluent Regulations MDMER
metal leaching ML
Metal Mining Effluent Regulations MMER
Methyl isobutyl carbinol MIBC
Metso Minerals Industries Inc. Metso
mineral titles online. MTO
Mining Rock Mass Rating MRMR
Ministry of Energy and Mines MEM
Ministry of Environment MOE
Ministry of Forests MOF
Ministry of Forests, Lands and Natural Resource Operations MFLNRO
Ministry of Forests, Lands and Natural Resource Operations Road Division MFLNRORD
Ministry of Transportation and Infrastructure MOTI
Mitchell Diversion Tunnel MDT
Mitchell Thrust Fault MTF
Mitchell Valley Drainage Tunnel MVDT
Mitchell-Treaty Twinned Tunnels MTT
molybdenum Mo
Moose Mountain Technical Services MMTS
motor control centre MCC
Multiple Pulse in Air MpiA
Municipal Wastewater Regulation MWR
Nass Timber Supply Area Nass TSA

 

Seabridge Gold Inc.xxiii221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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Nass Wildlife Area NWA
National Instrument 43-101 NI 43-101
nearest neighbor NN
Neil S. Seldon & Associates Ltd. NSA
net cash flow NCF
net present value NPV
net smelter price NSP
net smelter return NSR
Newcrest Mining Limited Newcrest
Newhawk Gold Mines Ltd. Newhawk
Newmont Exploration of Canada Ltd. Newmont
nickel Ni
Nisga’a Final Agreement NFA
Nisga’a Lisims Government NLG
non-potentially acid generating NPAG
Nordic Minesteel Technologies Inc. NMT
North American Datum NAD
North Pit Wall Dewatering Adit NPWDA
North Treaty Access Road NTAR
north N
Northwest Transmission Line NTL
Operator Interface Stations OIS
Ore Control System OCS
Ore Preparation Complex OPC
personal protective equipment PPE
Phelps Dodge Corp. Phelps Dodge
phosphorus P
pit slope angle PSA
Placer Dome Placer
Placer Dome Research Centre PDRC
Pocock Industrial Inc. Pocock
potassium amyl xanthate PAX
potentially acid generating PAG
Prefeasibility Study PFS
Preliminary Economic Assessment PEA
Pretium Resources Inc. Pretium
PricewaterhouseCoopers PwC
prilled ammonium AN Prill
probable maximum flood PMF
Process Tailing and Management Area PTMA
programmable logic controller PLC
Provincial Sales Tax PST
pyrite Py
Qualified Person QP
quality assurance QA

 

Seabridge Gold Inc.xxiv221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

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quality control QC
Quantitative Evaluation of Minerals by Scanning QEMSCAN®
quartz vein fragments QABX
Raewyn Copper RC
Raewyn fault RF
real-time kinematic RTK
regional study area RSA
regulation Reg.
request for information RFI
Resource Modelling Inc. RMI
reverse osmosis RO
Revised Statutes of British Columbia RSBC
rhenium Re
rock quality designation RQD
Rock Storage Facility RSF
run-of-mine ROM
Seabridge Gold Inc. Seabridge
selenium Se
semi-autogenous grinding mill comminution SMC
semi-autogenous grinding SAG
SGS Minerals Services SGS
silica SiO2
silver Ag
site water management SWM
S.J.V. Consultants Ltd. SJV
Skeena fold and thrust belt SFTB
Snowfields Slide Dewatering Adit SSDA
Society for Mining, Metallurgy, and Exploration SME
sodium cyanide NaCN
sodium hydrosulphide NaHS
sodium hydroxide NaOH
sodium silicate Na2SiO3
sodium sulphide Na2S
solids liquid separation SLS
south S
Special Use Permit SUP
Species at Risk Act SARA
Standard Penetration Test SPT
standard reference material SRM
sulphide S-2
sulphidization, acidification, recycling, and thickening of precipitate SART
sulphur S
Sulphurets Fault Zone SFZ
Sulphurets Thrust Fault STF
Sulphurets-Mitchell Conveyor Tunnel SMCT

 

Seabridge Gold Inc.xxv221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

sulphuric acid H2SO4
Surface Science Western SSW
Sustainable Resource Management Plan SRMP
System Impact Study SIS
Tailing Management Facility TMF
Tahltan Nation Development Corporation TNDC
Tariff Supplement TS
temporary water treatment plants TWTP
tennantite Ten
Tetra Tech, Inc. Tetra Tech
tetrahedrite Tt
The Claim Group Inc. TCG
the Environmental Assessment Application/Environmental Impact Statement the Application/EIS
time domain electromagnetic TDEM
total sulphur ST
total suspended sediments TSS
treatment charge/refining charge TC/RC
Treaty Creek Access Road TCAR
ultra-high frequency UHF
Universal Transverse Mercator UTM
University of British Columbia UBC
Unlimited Liability Corporation ULC
valued component VC
Voice over Internet Protocol VoIP
Water Storage Dam WSD
Water Storage Facility WSF
Water Treatment Plant WTP
weak acid dissociable WAD
west W
WN Brazier Associates Inc. Brazier
Wood Canada Limited. Wood
work breakdown structure WBS
work index Wi
x-ray fluorescence XRF
Z-Axis Tipper Electromagnetic ZTEM
zinc Zn

 

Seabridge Gold Inc.xxvi221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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1.0Summary

 

1.1Introduction

 

Seabridge Gold Inc.’s (Seabridge) Kerr-Sulphurets-Mitchell (KSM) Property (the Property) involves the development of major gold-copper-silver-molybdenum deposits located in northwest British Columbia (BC). KSM includes five major mineralized zones, identified as the Mitchell, East Mitchell, Kerr, Sulphurets, and Iron Cap deposits. KSM Mining ULC (KSMCo) is a wholly owned subsidiary of Seabridge Gold Inc.

 

KSMCo has received environmental assessment approvals and early-stage construction permits. Early construction works currently underway include the establishment of access roads, construction camps and power infrastructure.

 

In this report, “KSM Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” (the Report), the KSM mine development has been evaluated as a Prefeasibility Study (2022 PFS) using open pit mining of the Mitchell, East Mitchell and Sulphurets deposits. A Preliminary Economic Assessment (2022 PEA) is a stand-alone mine plan that has been undertaken to evaluate a potential future expansion of the KSM mine to the Iron Cap and Kerr deposits after the 2022 PFS mine plan has been completed. None of the Mineral Resources incorporated into the 2022 PEA mine plan have been used in the 2022 PFS mine plan.

 

This Report was prepared under the direction of Tetra Tech, Inc. (Tetra Tech), for Seabridge in 2022. The following independent consultants contributed to the Report:

 

Moose Mountain Technical Services (MMTS)

 

Wood Canada Limited (Wood)

 

Klohn Crippen Berger Ltd. (KCB)

 

ERM Consultants Canada Ltd. (ERM)

 

WN Brazier Associates Inc. (Brazier)

 

BGC Engineering Inc. (BGC)

 

WSP Golder Inc. (WSP Golder)

 

The results of the economic analyses for the 2022 PFS and 2022 PEA represent forward-looking information that is subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented in this Report. Forward-looking statements in this report include, but are not limited to, statements with respect to future metal prices, the estimation of Mineral Resources and Mineral Reserves, the estimated mine production and metals recovered, the estimated capital and operating costs, and the estimated cash flows generated from the planned mine production for the different development options. The material factors or assumptions used to develop the forward-looking information are identified in the relevant sections of this Report.

 

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KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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1.2Property Location and Ownership

 

The Property is located in the coastal mountains of northwest BC approximately 65 km by air north-northwest of Stewart, BC; and 21 km south-southeast of the former Eskay Creek Mine. Figure 1.1 is a general location map of the Property.

 

Figure 1.1 General Location Map

 

 

Source: Seabridge (2022)

 

The KSM Property comprises five discrete claim blocks and a group of placer claims. Placer claims are on only part of the westernmost claim block of the KSM Property. Claim blocks of the KSM Property are referred to as:

 

the KSM claims

 

the Seabee claims

 

the Tina claims

 

the Treaty Creek Switching Station claims

 

the East Mitchell (formerly named Snowfield) claim

 

Seabridge Gold Inc.1-2221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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The five KSM claim blocks include 80 mineral claims (cell and legacy) and 2 mining leases with a combined area of 42,052.40 ha. There are also 17 KSM placer claims held by KSM Mining ULC covering part of the KSM claims. The placer claims secure rights in a historically designated placer district. The Claim Group Inc. (TCG) acts as Agent on behalf of Seabridge with respect to maintaining all pertinent records associated with the KSM Property tenures. All claims and leases are in good standing under the Mining Tenure Act of BC and are recorded as owned 100% by KSMCo.

 

1.3Accessibility, Climate, Local Resources, Physiography, and Infrastructure

 

The Property lies within the rugged coastal mountains of northwestern BC, with elevations ranging from 520 masl in Sulphurets Creek Valley, to over 2,300 masl at the highest peaks. The climate is generally that of a temperate or northern coastal rainforest, with sub-arctic conditions at high elevations.

 

Construction is currently underway on KSM’s 30 km long Treaty Creek access road (TCAR) that will connect the KSM Process and Tailings Area (PTMA) to Highway 37, and the 33 km long Coulter Creek access road (CCAR) that will connect the mine area to Highway 37 via the 59 km long Eskay Creek mine resource road.

 

KSM will connect to BC Hydro’s existing Northwest Transmission Line (“NTL”) at BC Hydro’s Treaty Creek Switching Station (“TCT”). This TCT, located adjacent to the NTL and Highway 37, 18 km south of Bell 2 Lodge, is scheduled to be completed at the end of 2024. KSM Mining has completed its design for a 30 km long 287 kV transmission line to interconnect the TCT and the KSM plant site. This KSM transmission line is currently anticipated to start construction in 2023 with completion and commissioning planned for late 2024 to be ready for connection to the TCT.

 

There are multiple deep-water loading facilities for shipping bulk mineral concentrates located in the ice-free Port of Stewart, BC. Those port facilities are currently used by the Red Chris Mine. The nearest railway is the Canadian National Railroad (CNR) Yellowhead route, which is located approximately 220 km southeast of the Property. This line runs east-west, and can deliver concentrate to deep water ports near Prince Rupert and Vancouver, BC.

 

1.4History

 

The modern exploration history of the Property began in the 1960s, with brief programs conducted by Newmont Mining Corp. (Newmont), Granduc, Phelps Dodge, and the Meridian Syndicate. All of these programs were focused towards gold exploration. Various explorers were attracted to this area due to the numerous large, prominent pyritic gossans that are exposed in alpine areas. There is evidence that prospectors were active in the area prior to 1935. Several short hole, reconnaissance level drilling programs were undertaken between 1969 and 1991. The Sulphurets Zone was first drilled by Esso Minerals in 1969, Kerr was first drilled by Brinco Ltd. In 1985, Mitchell Creek by Newhawk in 1991, and Iron Cap by Esso Minerals in 1980.

 

Seabridge Gold Inc.1-3221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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In 1989, Placer Dome acquired a 100% interest in the Kerr deposit from Western Canadian Mines; in the following year, they acquired the adjacent Sulphurets Property from Newhawk. The Sulphurets Property also hosts the Mitchell Creek deposit and other mineral occurrences. In 2000, Seabridge acquired a 100% interest from Placer Dome in both the Kerr and Sulphurets properties, subject to capped royalties.

 

In 2020, Seabridge acquired the Snowfield deposit from Pretium Resources and renamed it to East Mitchell as it is geologically the sheared off top of the Mitchell deposit that was displaced eastward.

 

1.5Geological Setting and Mineralization

 

The Property lies within “Stikinia”, a long-lived volcanic island-arc terrane that extends over much of the Canadian Cordillera. It was accreted onto the Paleozoic basement of the North American continental margin in the Middle Jurassic. Early Jurassic sub-volcanic intrusive complexes in the Stikinia terrane host several large Cu-Au porphyry deposits including the KSM deposits.

 

The Kerr deposit is centered on a north-south trending, steep westerly dipping tabular intrusive complex with a strike extent of 2,400 m, a width of 800 m, and vertical extent of 2,200 m. Mineralization extends several ten’s of meters into the host sedimentary rocks. The Sulphurets deposit is composed of stacked thrust fault panels of Triassic and Jurassic volcano-sedimentary strata intruded by a number of dykes and stocks. It forms a lens dipping 30 degrees northwest extending 2,200 m horizontally, 550 m down dip, with a thickness of up to 330 m. The Mitchell Zone is underlain by intrusive, volcanic, and clastic rocks that are exposed in an erosional window below the shallow dipping Mitchell Thrust Fault (MTF). Mineralization is genetically and spatially related to the Early Jurassic Mitchell intrusive complex of diorite, monzodiorite, and granodiorite stocks and dykes. Mineralization also permeates into surrounding sedimentary and volcanic rocks, and in total extends 1,000 m east-west and 850 m north-south, with a vertical extent of 1,100 m. The Mitchell complex comprises three successive intrusive phases accompanied by the development of different hydrothermal assemblages, veining and mineralization. The East Mitchell deposit is the upper portion of the Mitchell deposit, displaced some 1.5 km to the southeast by the MTF during Cretaceous age compressive deformation that produced the regional Skeena Fold and Thrust belt. The Iron Cap deposit is also structurally above the MTF. It is a tabular body striking north-south and dipping 60 degrees to the west, extends 1,500 m along strike, 1,500 m down dip, and is 800 m in thickness.

 

The KSM deposits feature many characteristics typical of gold-enriched, diorite hosted calc-alkaline porphyry copper deposits, with gold, molybdenum, and silver at low concentrations, occurring as fine disseminations in quartz veinlet stockworks with accompanying pyrite, pervasively dispersed over hundreds of metres. All of the deposits are at least partially exposed at the surface, are largely unoxidized, and have had significant portions eroded away by glacial processes.

 

Seabridge Gold Inc.1-4221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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1.6Mineral Resources

 

The KSM Mineral Resource estimate summarized in Table 1.1 below is the current and only Mineral Resource estimate used in this Report.

 

Mineral Resources estimates for the Property were reviewed, validated and approved by Wood, with Henry Kim, P.Geo. taking responsibility for the estimates. The four mineralized zones, Kerr, Sulphurets, Mitchell and East Mitchell, and Iron Cap, were modeled separately. As more understanding was gained after each annual drilling campaign, individual block models were created for each area. Grade interpolation parameters have also evolved over time, reflecting changes required for modeling deeper mineralization intersected below the Kerr and Iron Cap deposits. A variety of basic descriptive statistics and spatial analyses were completed for each area upon the completion of annual drilling campaigns. These investigations include the generation of grade distribution tables, grade histograms, cumulative probability plots, grade box plots, grade contact plots, down-hole variograms, and directional variograms. In addition, new drill hole results were typically compared against the previous grade model to assess model performance.

 

The Mineral Resources for the various KSM mineralized zones are constrained within conceptual open pit and block cave mining shapes to support reasonable prospects for eventual economic extraction as outlined in the CIM Definition Standards for Mineral Resources and Mineral Reserves (CIM, 2014). The following gold, copper, silver, and molybdenum metal prices were used for determining block NSR values, US$1,300/oz, US $3.00/lb, US $20.00/oz, and US $9.70/lb, respectively. Open pit and underground mining costs of Cdn$1.80 to 2.20/tonne and Cdn$6.00 to Cdn$7.00/tonne were used to establish conceptual open pit and underground resource shapes, along with a processing and G&A cost of Cdn$9.00/tonne for Kerr, Sulphurets and Iron Cap, and Cdn$10.75 to Cdn$11.20/tonne for Mitchell and East Mitchell.

 

The conceptual open pit and underground mining shapes were generated for each resource area based on calculated block model NSR values. The NSR values were generated for each deposit. MMTS generated conceptual pits for the Kerr, Sulphurets, Mitchell and East Mitchell deposits using MineSight® software and Lerchs-Grossmann algorithms. WSP Golder developed conceptual block cave footprints for Kerr and Iron Cap using the block NSR values and Geovia’s PCBC Footprint Finder software. The footprint polygons were extruded vertically based on guidance from WSP Golder.

 

The draw point extraction elevations were extruded vertically to create 3D solids that were used for resource tabulation. Conceptual caves were clipped against surface topography (Iron Cap) or conceptual resource pit (Kerr). Mineral Resources are determined, at Cdn$10.75 to Cdn$11.25 and Cdn$16 NSR cutoffs for open-pit constrained and underground mining constrained Mineral Resources, respectively.

 

Seabridge Gold Inc.1-5221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

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Table 1.1 summarizes the estimated Measured, Indicated, and Inferred Mineral Resources for each zone.

 

Table 1.1KSM Mineral Resources

 

Measured Mineral Resources

 

 

Project

Cut Off

(Cdn$/t)

Tonnes (000) Gold Copper Silver Molybdenum
Grade
(g/t)
Ounces (000) Grade
(%)
Pounds (millions) Grade
(g/t)
Ounces (000) Grade (ppm) Pounds (millions)
KSM: NSR:                  
Mitchell $10.75 691,700 0.68 15,124 0.19 2,876 3.3 72,831 52 79
East Mitchell $11.25 1,012,800 0.65 21,098 0.11 2,514 1.8 59,233 89 198
KSM Total   1,704,500 0.66 36,222 0.14 5,390 2.4 132,064 74 277

 

Indicated Mineral Resources

 

 

Project

Cut Off
(Cdn$/t)
Tonnes (000) Gold Copper Silver Molybdenum
Grade
(g/t)
Ounces (000) Grade
(%)
Pounds (millions) Grade
(g/t)
Ounces (000) Grade (ppm) Pounds (millions)
KSM:

$10.75-$11.25 NSR Pits

 

$16 NSR

UG

                 
Mitchell 1,667,000 0.48 25,935 0.14 5,120 2.8 149,160 66 241
East Mitchell 746,200 0.42 10,080 0.08 1,390 1.7 41,814 79 130
Sulphurets 446,000 0.55 7,887 0.21 2,064 1.0 14,339 53 52
Kerr 374,000 0.22 2,660 0.41 3,405 1.1 13,744 5 4
Iron Cap 423,000 0.41 5,576 0.22 2,051 4.6 62,559 41 38
KSM Total 3,656,200 0.44 52,138 0.17 14,030 2.4 281,616 58 465

 

Measured plus Indicated Mineral Resources

 

 

Project

Cut Off
(Cdn$/t)
Tonnes (000) Gold Copper Silver Molybdenum
Grade
(g/t)
Ounces (000) Grade
(%)
Pounds (millions) Grade
(g/t)
Ounces (000) Grade (ppm) Pounds (millions)
KSM:

$10.75-$11.25 NSR Pits

 

$16 NSR

UG

                 
Mitchell 2,358,700 0.54 41,059 0.15 7,996 2.9 221,991 62 320
East Mitchell 1,759,000 0.55 31,178 0.10 3,904 1.8 101,047 85 328
Sulphurets 446,000 0.55 7,887 0.21 2,064 1.0 14,339 53 52
Kerr 370,000 0.22 2,660 0.41 3,405 1.1 13,744 5 4
Iron Cap 423,000 0.41 5,576 0.22 2,051 4.6 62,559 41 38
KSM Total 5,356,700 0.51 88,360 0.16 19,420 2.4 413,680 63 742

 

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Inferred Mineral Resources

 

 

Project

Cut Off
(Cdn$/t)
Tonnes (000) Gold Copper Silver Molybdenum
Grade
(g/t)
Ounces (000) Grade
(%)
Pounds (millions) Grade
(g/t)
Ounces (000) Grade (ppm) Pounds (millions)
KSM:

$10.75 NSR Pits

 

$16 NSR

UG

                 
Mitchell 1,282,600 0.29 11,819 0.14 3,832 2.5 102,228 47 133
East Mitchell 281,100 0.37 3,372 0.07 403 2.3 21,112 61 38
Sulphurets 223,000 0.44 3,155 0.13 639 1.3 9,320 30 15
Kerr 1,999,000 0.31 19,823 0.40 17,720 1.8 114,431 23 103
Iron Cap 1,899,000 0.45 27,474 0.30 12,556 2.6 158,741 30 126
KSM Total 5,684,700 0.36 65,643 0.28 35,150 2.2 405,832 33 415

 Notes:

1.The effective date for the Mineral Resource Estimate for Mitchell and East Mitchell is March 31, 2022, and for Kerr, Sulphurets and Iron Cap is December 31, 2019.

2.The Mineral Resource estimates have been reviewed and approved by Henry Kim P.Geo., an independent Qualified Person. Mr. Kim verified the databases supporting the Mineral Resource estimates and conducted a personal inspection of the property and reviewed drill core from a range of representative drill holes at site and at the core storage facilities in Stewart, B.C. with Seabridge geology staff.

3.Mineral Resources were prepared in accordance with CIM Definition Standards for Mineral Resources and Mineral Reserves (May 10, 2014) and CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines (Nov 29, 2019).

4.Mineral Resources were constrained within mineable shapes depending on their mining methods.

5.Mineral Resources are reported inclusive of those Mineral Resources that were converted to Mineral Reserves. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

6.Following metal prices were used to determine Mineral Resources: US$1300/oz Au, US$3/lb Cu, US$20/oz Ag, and US$9.7/lb Mo.

7.Details regarding key assumptions, parameters, and methods used in estimating the Mineral Resources are included in Section 14.0 of this Report.

8.Numbers may not add due to rounding.

 

1.7Mining Methods

 

The 2022 PFS uses conventional large-scale open pit mining methods. Pit phases at the Mitchell, East Mitchell, and Sulphurets deposits have been engineered based on the results of economic pit limit analysis. Pit size is limited by the available permitted tailings storage volume of 2.29 Bt. Starter pits have been selected in higher-grade areas.

 

1.7.1Mineral Reserve Estimate

 

Mineral Reserves for the 2022 PFS are based on open pit mining of the Mitchell, East Mitchell and Sulphurets deposits. Waste to ore cut-offs were determined using an NSR for each block in the model. NSR is calculated using prices and process recoveries for each metal accounting for all off-site losses, transportation, smelting and refining charges. Metal prices of US$1,300/oz gold, US$3.00/lb copper, US$20.00/oz silver and US$9.70/lb molybdenum and a foreign exchange rate of US$0.79 to Cdn$1.00 have been used in the NSR calculations.

 

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Lerchs-Grossman (“LG”) pit shell optimizations were used to define open pit mine pit limits in the 2022 PFS. Production is limited by the permitted tailings volume of 2.29 Bt. Open pit designed phases use updated geotechnical studies based on most recent site investigation programs.

 

Mineral reserves have been estimated using the updated pit designs. The open pit minimum NSR cut-off grade is based on an estimated process operating cost of Cdn$11.00/t. Process operating costs include plant processing (including crushing/ore transport costs where applicable), G&A, surface service, tailing construction, and water treatment costs. A premium cut-off grade of Cdn$25.00/t is used until the end of Year 5 to maximize the net present value (NPV) and minimize the time to payback of initial capital.

 

Proven and Probable Mineral Reserves for the KSM mineral deposits are stated in Table 1.2.

 

Table 1.2KSM Proven and Probable Mineral Reserves

 

Reserve
Category
Deposit Ore
(Mt)
Diluted Grades Contained Metal
Au
(g/t)
Cu
(%)
Ag
(g/t)
Mo
(ppm)
Au
(Moz)
Cu
(Mlb)
Ag
(Moz)
Mo
(Mlb)
Proven Mitchell 483 0.74 0.20 3.3 49 11.5 2,161 51 53
East Mitchell 814 0.69 0.11 1.8 91 18.1 2,043 47 163
Sulphurets 0 0.00 0.00 0.0 0 0.0 0 0 0
Total Proven 1,297 0.71 0.15 2.4 75 29.6 4,203 98 215
Probable Mitchell 452 0.59 0.15 2.5 74 8.6 1,458 36 74
East Mitchell 392 0.46 0.09 1.7 84 5.8 784 21 73
Sulphurets 151 0.68 0.26 1.0 70 3.3 874 5 23
Total Probable 995 0.55 0.14 1.9 77 17.7 3,116 62 170
Proven + Probable Mitchell 935 0.67 0.18 2.9 61 20.1 3,619 87 126
East Mitchell 1,206 0.62 0.11 1.8 89 23.9 2,826 68 236
Sulphurets 151 0.68 0.26 1.0 70 3.3 874 5 23
Total Proven + Probable 2,292 0.64 0.14 2.2 76 47.3 7,320 160 385

Notes:

1.The Mineral Reserve estimates were reviewed by Jim Gray, P.Eng. (who is also the independent Qualified Person for these Mineral Reserve estimates), reported in accordance with 2014 CIM Definition Standards and 2019 CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines, and have an effective date of May 26, 2022.
2.Mineral Reserves are based on the 2022 PFS all open pit Life of Mine plan.
3.Mineral Reserves are mined tonnes and grade, the reference point is the mill feed at the primary crusher and includes consideration for operational modifying factors.
4.Mineral Reserves are reported at NSR cut-offs that vary between of Cdn$11/t and Cdn$25/t using the following assumptions: metal prices of US$1300/oz Au, US$3.00/lb Cu, US$20/oz Ag, and US$ 9.70/lb Mo at a currency exchange rate of 0.79 US$ per 1.00 Cdn$; Copper concentrate terms are 96% payable Cu; 97.8% payable Au; 90% payable Ag, molybdenum concentrate terms are 99% payable. Offsite costs (smelting, refining, transport, and insurance) are Cdn$281 per tonne of copper concentrate and Cdn$5,527 per tonne of molybdenum concentrate; doré terms are US$2/oz offsite costs (refining, transport and insurance), 99.8% Au payable, and 90% Ag payable; metallurgical recovery projections vary depending on metallurgical domain and metal grades and are based on metallurgical test work.
5.The NSR cut-off is varied from Cdn$11/t to Cdn$25/t and covers the estimated process operating cost of Cdn$10/t for ore processing, G&A, surface service, tailings, and water treatment costs.
6.Mineral Reserves account for mining loss and dilution.
7.Mineral Reserves are a subset of the mineral resource.
8.Numbers have been rounded as required by reporting guidelines.

 

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1.7.2Mine Production Plan

 

All ore from the 33 years of mine life is derived from open pits. Preproduction in Years -2 and Year -1 mine waste from the top benches of the Mitchell starter pit to expose ore at the 1080 m elevation. Mill feed ramps up to a nominal 130,000 t/d by Year 2 followed by a 50% processing plant expansion by Year 3 to a nominal 195,000 t/d for the remaining mine life. Life of mine (LOM) production is summarized in Figure 1.2.

 

Ore is mined primarily from the Mitchell open pit from Years 1 to 7. A small high grade pit in the East Mitchell upper zone is blended with Mitchell ore from Years 3 to 5. A small Sulphurets pit is mined from Years 6 to 12. The main East Mitchell pit starts ramping up from Year 8 and becomes the primary mill feed source from Year 12 forward. Final Mitchell pit phases are mined from Years 18 to 26. Final East Mitchell pit phase is mined from Years 24 to Year 33. A portion of waste from the final East Mitchell pit phase is backfilled into the mined-out Mitchell pit.

 

Figure 1.22022 PFS Mill Feed Production Schedule

 

Source: MMTS (2022)

 

1.8Mineral Processing and Metallurgical Testing

 

Several wide-ranging metallurgical test programs have been carried out since 2007 to assess the metallurgical responses of the mineral samples from the KSM deposits, especially the samples from the Mitchell deposit.

 

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The primary economic metals at KSM are gold and copper. Primary copper bearing mineral is chalcopyrite. Gold and silver are associated primarily with chalcopyrite and pyrite. There is variability in the pyrite to chalcopyrite ratio between the deposits.

 

The test results indicate that the mineral samples from the KSM deposits are amenable to the proposed KSM flowsheet including:

 

copper-gold-silver-molybdenum bulk rougher flotation followed by gold--silver bearing pyrite flotation

 

regrinding the bulk rougher concentrate followed by three stages of cleaner flotation to produce a copper-gold-silver-molybdenum bulk cleaner flotation concentrate

 

molybdenum separation of the bulk cleaner flotation concentrate to produce a molybdenum concentrate and a copper-gold concentrate containing associated silver

 

cyanide leaching of the gold- silver bearing pyrite flotation concentrate and the scavenger cleaner tailing to further recover gold and silver values as doré, the cyanide circuit includes SART and AVR processes to recover weak acid dissociable cyanide for reuse and dissolved copper for sale with copper concentrate

 

the flotation tailing will be sent to the flotation tailing storage cells with the TMF; the leach residue will be destructed for residual cyanide prior to being sent to the lined CIL Residue Cell, supernatants from both the cells will be reclaimed separately to the process plant for reuse as process makeup water.

 

Primary crushing is by gyratory crushers at the Mitchell ore processing complex (Mitchell OPC). Coarse ore from the primary crushers is transported by train through the Mitchell-Treaty Twinned Tunnel (MTT) to the processing facility in the PTMA. The MTT will also be used for electrical power transmission and the transport of personnel and supplies for mine area operations.

 

Coarse ore from the train is conveyed to two coarse ore stockpiles, followed by secondary cone crushing and tertiary crushing by HPGR. Fine ore from the HPGR is fed to ball mills followed by copper-gold-silver-molybdenum bulk rougher flotation and pyrite flotation of copper rougher tailings. Bulk flotation concentrates are reground prior to cleaning flotation that produces bulk copper-gold-silver flotation concentrate and molybdenum concentrate products and a gold-silver bearing pyritic product for cyanide leaching. Final products include a copper-gold-silver concentrate, gold-silver doré, and a molybdenum concentrate.

 

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1.92022 PFS Infrastructure

 

1.9.1Geohazards

 

Geohazard and risk assessments were completed for the proposed facilities within the KSM footprint. As expected for a mountainous, high-relief property site, snow avalanche and landslide hazards exist, with the potential to affect mine construction, operations, and closure.

 

BGC reviewed the proposed mine area and surrounding terrain for potential geohazards, including the identification of snow avalanche paths and potential landslides, utilizing aerial photographs and satellite imagery. BGC completed ground-truthing of potential geohazards; the preliminary design of mitigation structures were completed by those responsible for the various facilities at risk from the identified geohazards.

 

Geohazard scenarios were identified for the mine facilities. Using unmitigated geohazard levels as a baseline, these scenarios were assessed in terms of risk to human safety, economic loss, environmental loss, and reputation loss. Geohazard risk levels were assigned to each scenario with ratings ranging from very low to very high.

 

Mitigation strategies have been identified to reduce the high and very high risk scenarios to a target residual risk not exceeding moderate. Further risk reduction will be achieved where practical and cost-efficient and as part of the detailed design of specific facilities.

 

1.9.2Tailings Management

 

The TMF for the 2022 PFS will be constructed in three cells: the North and South cells for flotation tailings, and a lined carbon-in-leach (CIL) Residue Cell for CIL residue tailings. The cells are confined between four dams (North Dam, Splitter Dam, Saddle Dam, and Southeast Dam) located within the Teigen-Treaty Creek cross-valley. Design criteria for the dams are based on the Canadian Dam Association (CDA) guidelines.

 

De-pyritized flotation tailings will be stored in the North and South cells. The pyrite bearing CIL residue tailings will be stored in the lined CIL Residue Cell. In total, the TMF will have a capacity of 2.29 Bt as described in the permitted Certified Project Description.

 

An Independent Geotechnical Review Board (IGRB) was established in January 2015 to independently review and to provide expert oversight, opinion, and advice to Seabridge on the design, construction, operational management, and ultimate closure of the TMF and Water Storage Dam (WSD). The IGRB reviews the TMF and WSD data collection and design processes on an ongoing basis to ensure that these structures meet internationally accepted standards and practices which effectively minimize risks to employees, lands, and communities.

 

A Best Available Tailings Technology (BATT) review in August 2015 concluded that the existing TMF design, consisting of centerline dams constructed with double cycloned sand and a till core in association with hydraulic tailings deposition, is the best available technology for tailings deposition, and the most environmentally responsible design to minimize long-term risks associated with the proposed TMF. This conclusion confirms the findings from KSM’s IGRB that the TMF’s design is robust and appropriate for KSM’s site-specific characteristics.

 

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1.9.3Mine Site Water Management

 

The overall site water management strategy for the 2022 PFS was reviewed and approved during the EA review process.

 

Diversion Tunnels and Surface Diversion

 

Two main diversion tunnel routes will be required to route non-contact water from the Mitchell and McTagg valleys around the mine site.

 

Lined surface diversion channels will be constructed progressively during operations, along the contact of the RSF and the hillside, to divert surface flows.

 

Water Storage Facility

 

All contact water from the mine site areas (open pits, RSFs, roads, infrastructure) will be directed to the WSF, located in the lower Mitchell Creek area. The WSF will be formed with a 165 m high rock fill asphalt core dam built to full height by Year -1 and is sized to store annual freshet flows and volumes resulting from a 200-year wet year. The core zones of the WSF dam will be founded on competent sedimentary rock foundations. Seepage will be controlled by the asphalt core in the dam and the dam foundation will be grouted. A seepage collection pond will collect seepage water beyond the toe of the main dam and return it to the WSF.

 

Water Treatment

 

Mine area contact water will be treated with a High Density Sludge (HDS) lime water treatment plant (WTP), of which the discharge from the plant was approved as a component of the EA review process. A Selenium WTP will be constructed and operational by Year 5 to treat up to 500 L/s of seepage principally from the RSF and select point sources within Mitchell Valley with selenium loaded waters, compared to lower concentrations within the WSF.

 

The HDS WTP and the WSF will be operational before mill start-up to allow pre-production activity in the Mitchell Valley and Mitchell pit area.

 

1.9.4Tunneling

 

Three major tunnels will be excavated during the construction period:

 

Mitchell Treaty Twinned Tunnels (MTT)

 

Mitchell Diversion Tunnel (MDT)

 

McTagg Diversion Tunnel (MTDT)

 

These tunnels are classified as either infrastructure tunnels (MTT) or water tunnels (MDT and MTDT). Additional tunnels will be constructed at various times during mine operations for diversion of contact water around mine facilities in Mitchell Valley or non-contact water around the east side of the TMF.

 

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The tunnels will be constructed in accordance with BC Mines Act and Health, Safety and Reclamation Code for Mines in BC using conventional drill and blast techniques and will follow the conditions contained within the License of Occupation for the MTT issued in September 2014 by the Government of BC.

 

1.9.5Mine To Mill Ore Transport System

 

Ore will be crushed at the Mitchell OPC, loaded onto trains and transported to the Treaty coarse ore stockpile (COS) via the 22.9 km twin MTT. The trains are autonomous and controlled from a control centre at Treaty. The electric drives rely on regenerative braking which is input back into the grid.

 

The ore transport is configured to start operations at a nominal production rate of 130,000 tpd and ramps up to 195,000 tpd for the start of Year 3.

 

The train system will also handle personnel, freight, and fuel requirements between Treaty and Mitchell. The power cable for the mine will also be in the MTT

 

1.10Environmental Studies, Permitting, and Social or Community Impact

 

The KSM mine development plan was subject to the BC Environmental Assessment Act (BCEAA, the Act), the Canadian Environmental Assessment Act- 1992 (CEAA), and Chapter 10 of the Nisga’a Final Agreement (NFA). In this section of the Report, the term “Project” refers to a mine development plan that would substantially support what is contemplated in the 2022 PFS.

 

KSM has successfully gone through the provincial and federal environmental assessment review processes, and the appropriate certificates/approvals have been obtained. Additionally, permits for early-stage construction activities, continuation of exploration, and certain permit and project approval renewals have also been obtained. Seabridge continues to advance permitting to allow for the construction of the 2022 PFS, as well as to expand exploration activities.

 

1.10.1Benefit Agreements Tahltan Nation

 

On July 8, 2019, the Tahltan Nation and KSMCo announced that the parties reached agreement on the terms of a Co-operation and Benefits Agreement in connection with the Project (IBA). The IBA provides a framework for the parties to work together in relation to the Project and includes commitments to economic benefits and environmental management of the land.

 

Nisga’a Nation

 

On June 16, 2014, KSMCo entered into a comprehensive Benefits Agreement with the Nisga’a Nation in respect of the Project. The Benefits Agreement establishes a long-term co-operative working relationship between KSMCo and the Nisga’a Nation, under which the Nisga’a Nation will support the mine development and participate in economic benefits from the Project Environmental Settings and Studies.

 

1.10.2Closure and Reclamation

 

In the 2022 PFS, the KSM Mine will be closed in accordance with the closure plan outlined in Section 20.6, and in further detail in the Application/EIS (Rescan 2013).

 

The conceptual closure and reclamation plan has three objectives that provide assurance to the Province that the site will be left in a condition that will limit any future liability to the people of BC:

 

to provide stable landforms

 

to re-establish productive land use

 

to protect terrestrial and aquatic resources.

 

1.112022 PFS Capital Cost Estimate

 

An initial capital of US$6.432 billion is estimated for the 2022 PFS. All currencies in are expressed in US dollars, unless otherwise stated. Costs have been converted using a fixed currency exchange rate of US$0.77 to Cdn$1.00. The expected accuracy range of the capital cost estimate is +25%/-10%.

 

A summary of the 2022 PFS initial capital costs is shown in Table 1.3.

 

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Table 1.32022 PFS Initial Capital Cost Summary

 

Major
Area
No.
Major Area Description Cost
(US$ M)
1 – Direct Costs
1.1 Mine Site 1,420
1.2 Process 2,003
1.3 TMF 513
1.4 Environmental 15
1.5 On-site Infrastructure 39
1.6 Off-site Infrastructure 76
1.7 Permanent Electrical Power Supply and Energy Recovery 121
Total Direct Costs 4,188
2 – Indirect Costs
2.91 Construction Indirect Costs 565
2.92 Spares 55
2.93 Initial Fills 26
2.94 Freight and Logistics 110
2.95 Commissioning and Start-up 7
2.96 EPCM 299
2.97 Vendor’s Assistance 29
Total Indirect Costs 1,090
3 – Owner’s Costs
3.98 Owner’s Costs 204
4 – Contingency
4.99 Contingency 949
2022 PFS Capital Cost Total 6,432

Notes: Costs have been rounded to the nearest million dollars.

 

Capital costs exclude reclamation and closure costs that are accounted for in the economic analysis. Sustaining capital costs were also estimated leveraging the same basis of information applied to the initial capital estimate with respect to vendor quotations, labour, and material costs. The sustaining capital costs total US$3.210 billion and consist of:

 

open pit mine development, principally mobile fleet replacement

 

process plant expansion

 

TMF expansions, mainly comprising dam raises and CIL basin expansions

 

indirect costs, including construction indirects, spares, freight, and logistics, EPCM, vendor assistance, and contingency.

 

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1.122022 PFS Operating Cost Estimate

 

The average operating cost for the 2022 PFS is estimated at US$11.36/t milled. The cost estimates are based upon budget prices in Q1/Q2 2022 or data from the database of the consulting firms involved in the cost estimates. The expected accuracy range of the operating cost estimate is +25%/-10%.

 

Average life of mine unit operating costs by areas are shown in Table 1.4:

 

Table 1.42022 PFS LOM Average Unit Operating Costs

 

Area Cost (US$/t milled)
Mining 3.31
Process 6.31
G&A and Site Services 1.06
Others 0.67
Total Operating Cost 11.36

Notes:1. Mining operating cost excludes pre-production cost of both open pits.
2.Others includes tailings management, water management, annual energy recovery credit and Provincial Sales Tax (PST).
3.Numbers may not add due to rounding.

 

1.132022 PFS Economic Evaluation

 

Tetra Tech prepared a Base Case economic evaluation for the 2022 PFS incorporating historical three-year trailing averages for gold, copper and silver metal prices of as of June 20, 2022. This approach is used because it is consistent with the 2016 PFS Base Case. Molybdenum price is based on a recent study for a primary molybdenum project. Two alternate cases are also presented: (i) The Recent Spot Case incorporating recent spot prices for gold, copper, silver and the US$/Cdn$ exchange rate; and, (ii) The Alternate Case that incorporates lower metal prices than used in the Base Case to demonstrate the 2022 PFS’s sensitivity to lower metal prices. The pre-tax and post-tax estimated economic results in U.S. dollars for all three cases are shown in Table 1.5 and Table 1.6, respectively.

 

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Table 1.52022 PFS Summary of the Pre-tax Economic Evaluations

 

  Unit 2022 PFS
Base
Case
2022 PFS
Spot
Case
2022 PFS
Alternate
Case
Gold US$/oz 1,742.00 1,850.00 1,500.00
Copper US$/lb 3.53 4.25 3.00
Silver US$/oz 21.90 22.00 20.00
Molybdenum US$/lb 18.00 18.00 18.00
Exchange Rate US$:Cdn$ 0.77 0.77 0.77
Undiscounted NCF US$ million 38,636 46,070 27,854
NPV (at 3%) US$ million 20,210 24,357 14,210
NPV (at 5%) US$ million 13,454 16,403 9,194
NPV (at 8%) US$ million 7,420 9,294 4,717
IRR % 20.1 22.4 16.5
Payback years 3.4 3.1 4.1
Cash Cost/oz Au US$/oz 275 164 351
Total Cost/oz Au US$/oz 601 490 677

Note: Net cash flow (NCF)

 

Table 1.62022 PFS Summary of the Post-tax Economic Evaluations

 

  Unit 2022 PFS
Base
Case
2022 PFS
Spot
Case
2022 PFS
Alternate
Case
Gold US$/oz 1,742.00 1,850.00 1,500.00
Copper US$/lb 3.53 4.25 3.00
Silver US$/oz 21.90 22.00 20.00
Molybdenum US$/lb 18.00 18.00 18.00
Exchange Rate US$:Cdn$ 0.77 0.77 0.77
Undiscounted NCF US$ million 23,933 28,630 17,133
NPV (at 3%) US$ million 12,264 14,889 8,467
NPV (at 5%) US$ million 7,944 9,814 5,238
NPV (at 8%) US$ million 4,061 5,254 2,332
IRR % 16.1 18.0 13.1
Payback years 3.7 3.4 4.3

Notes:

1.Operating and total cost per ounce of gold are after copper, silver and molybdenum credits.
2.Total cost per ounce includes all start-up capital, sustaining capital and reclamation/closure costs.
3.Results include consideration of Royalties and Impact Benefit Agreements payments.
4.The post-tax results include the B.C. Mineral Tax and provincial and federal corporate taxes.

 

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1.13.1Sensitivity Analysis

 

Tetra Tech investigated the sensitivity of NPV and IRR to the key variables. Using the 2022 PFS Base Case as a reference, each of key variables was changed between -30% and +30% in 10% increments while holding the other variables constant.

 

Sensitivity analyses were carried out on the following key variables:

 

gold, copper, silver, and molybdenum metal prices

 

exchange rate

 

capital costs

 

operating costs.

 

The analyses are presented graphically as financial outcomes in terms of post-tax NPV, and IRR. The NPV is most sensitive to gold price and exchange rate, followed by operating costs, copper price, and capital costs. The IRR is most sensitive to exchange rate, capital costs, and gold price, followed by copper price and operating costs. In general, sensitivity to metal price is roughly equivalent to sensitivity to metal grade. Financial outcomes are relatively insensitive to silver and molybdenum prices. The NPV and IRR sensitivities are presented in Figure 1.3 and Figure 1.4, respectively.

 

Figure 1.32022 PFS Sensitivity Analysis of Post-tax NPV at a 5% Discount Rate

 

Source: Tetra Tech (2022)

 

Figure 1.42022 PFS Sensitivity Analysis of Post-tax IRR

 

 

Source: Tetra Tech (2022)

 

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1.142022 Preliminary Economic Assessment

 

1.14.1Introduction

 

The 2022 PEA is a stand-alone mine plan that has been undertaken to evaluate a potential future expansion of the KSM mine to the Iron Cap and Kerr deposits after the 2022 PFS mine plan has been completed. None of the Mineral Resources incorporated into the 2022 PEA mine plan have been used in the 2022 PFS mine plan.

 

The 2022 PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the results of the 2022 PEA will be realized. Mineral Resources in the 2022 PEA mine plan are not Mineral Reserves and do not have demonstrated economic viability.

 

1.14.2Mining Methods

 

The 2022 PEA utilizes Measured, Indicated, and Inferred Mineral Resources in mine planning. Mill feed in the 2022 PEA is drawn from open pit mining of the Kerr deposits and underground block cave mining of the Iron Cap and Kerr deposits.

 

Approximately 93% of the mill feed would come from underground mine operations. The Kerr open pit will supplement the mill feed during the ramp up of Iron Cap underground mine. The Kerr underground mine ramp up will begin after the Iron Cap underground mine ramp up is completed and the Kerr open pit operations lapsed.

 

1.14.3Open Pit Mining

 

The 2022 PEA open pit mining is a conventional truck-shovel operation. pit design for the Kerr deposit includes a single mining phase. The Mineral Resources within the 2022 PEA mine plan is a subset of the Mineral Resources and quantities that are included in Section 14.0. The NSR cutoff for the PEA open pit is Cdn$10.75/t.

 

1.14.4Underground Mining

 

The 2022 PEA underground block caving mine designs for Iron Cap and Kerr are based on modeling using GEOVIA’s Footprint Finder software. Iron Cap has been designed for using battery-electric loaders and electric trains for material handling and employs both tele-operation and automation for these units. This enables a reduction to the number of primary ventilation intake and exhaust drifts due to less ventilation demand and operating cost savings from lower diesel fuel consumption, labour and equipment maintenance.

 

Iron Cap underground mining is estimated to have a production ramp-up period of six years, steady state production at 33 Mt/a for 17 years, and then ramp-down production for another six years. Kerr underground mining is estimated to have a production ramp-up period of five years, steady state production at 29 Mt/a for 20 total years with a seven year production dip during years where the operation transitions from the first to second lift. The underground pre-production period is four years for Iron Cap and five years for Kerr.

 

Underground mill feed production from Iron Cap starts in PEA Year 1 and Kerr in PEA Year 7. Indicated and Inferred Mineral Resources are included in the production plan.

 

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1.14.5Mine Production Plan

 

After initial ramp-up, the mill throughput averages 170,000 t/d until PEA Year 23 when Iron Cap mining is completed. Then the production rate is reduced to around 80,000 t/d as mill feed is provided from only from the Kerr underground mine. The 2022 PEA mine life is estimated at 39 years.

 

Table 1.7 and Figure 1.5 show a summary of the production tonnes and grades in the 2022 PEA mine plan.

 

Table 1.7Production Tonnes and Grade in 2022 PEA Mine Plan

 

Zone Mining Method Classifi-cation Tonnes (millions) Average Grades Contained Metal
Gold Copper Silver Gold Copper Silver
(g/t) (%) (g/t) M oz’s M lbs M oz’s
Iron Cap Block Cave M+I 58 0.62 0.28 3.2 1.1 354 5.9
Inferred 685 0.58 0.36 3.0 12.7 5,424 65.4
Kerr Open Pit M+I 117 0.26 0.51 1.4 1.0 1,315 5
Inferred 7 0.74 0.09 1.5 0.2 14 0
Block Cave M+I 48 0.25 0.53 1.3 0.4 557 2
Inferred 777 0.31 0.49 1.7 7.8 8,339 43.6
Total Mill Feed Mined M+I 223 0.35 0.45 1.8 2.5 2,226 13
Inferred 1,469 0.44 0.43 2.3 20.7 13,777 109
Note:The 2022 PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the results of the 2022 PEA will be realized. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

 

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Figure 1.52022 PEA Mill Feed Production Schedule

 

 

1.14.6Recovery Methods

 

The 2022 PEA process plant will have a peak process rate of 170,000 t/d. The plant will receive mill feed from Iron Cap and Kerr deposits.

 

The 2022 PEA process circuit incorporates three-stage crushing (tertiary crushing by HPGR), ball mill grinding, and flotation for the recovery of copper, gold, silver, and molybdenum. Leaching of gold-silver bearing pyritic products from Iron Cap and Kerr mineralization is excluded from the 2022 PEA.

 

1.14.72022 PEA Capital and Operating Costs

 

Initial capital costs of the 2022 PEA are estimated at US$1.5 billion. Sustaining capital costs over the 39-year mine life are estimated at US$12.8 billion and is dominated by capitalizing the underground mine development at Iron Cap and Kerr block caves. Initial capital and sustaining capital estimates for the 2022 PEA are summarized in Table 1.8.

 

Table 1.82022 PEA Capital Cost Estimate Summary

 

Area Initial Sustaining Total
US$ M US$ M US$ M
Direct Costs      
Mine 828 6,678 7,506
Process 0 651 651
TMF 74 664 738
On-site Infrastructure 26 573 599
Power Supply/Energy Recovery 0 112 112
Total Direct Capital 927 8,678 9,606
Indirect cost 253 1,249 1,502
Contingency 320 2,824 3,145
Total Estimated Capital Cost 1,500 12,752 14,252
Notes:1. Sums may not add due to rounding,

2.PST not included in table please see section 24.22 for details.
3.All closure costs are addressed separately in financial analyses outside of capital.
4.Most of the sustaining on-site infrastructure is included in the mine category.
5.Mine direct costs include for crushing, conveying and power infrastructure.

 

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Average mining, process, and G&A operating costs over the 2022 PEA calculated mine life (including waste mining and on-site power credits, excluding off-site shipping and smelting costs) are estimated at US$11.98/t milled (before base metal credits). A breakdown of estimated unit operating costs is shown in Table 1.9.

 

Table 1.92022 PEA LOM Average Unit Operating Costs

 

Area Cost
(US$/t milled)
Mining 4.99
Process 4.31
G&A and Site Services 1.89
Others 0.79
Total Operating Costs 11.98
Notes:1. Mining operating cost excludes pre-production cost of both open pit and underground mining.
2.Others includes tailings management, water management, annual energy recovery credit and PST.

 

1.14.8Economic Analysis

 

The 2022 PEA has been evaluated using a discounted cash flow (DCF) analysis. Cash inflows consist of annual revenue projections for the mine. Cash outflows such as capital (including pre-production costs), operating costs, taxes, and royalties are subtracted from the inflows to arrive at the annual cash flow projections. Cash flows are taken to occur at the end of each period.

 

Preproduction development period for the PEA was assumed to be 4 years, economic results were reported at the start of that 4 years period. Under the assumptions presented in this report, the 2022 PEA demonstrates positive economics. In the 2022 PEA Base Case, the post-tax NPV at a 5% discount rate is US$5.8 billion with an post -tax IRR of 18.9%. The post -tax payback of the initial capital investment is estimated to occur in 6.2 years after the start of production. The 2022 PEA Economic Results are presented in Table 1.10.

 

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Table 1.102022 PEA Economic Results

 

Financial Results 2022 PEA
Base Case
3yr Avg.
2022 PEA
Alternative
Case

2022 PEA
Recent Spot

Upside Case

Metal Prices:      
Gold (US$/oz) 1,742.00 1,500.00 1,850.00
Copper (US$/lb) 3.53 3.00 4.25
Silver (US$/oz) 21.90 20.00 22.00
Molybdenum (US$/lb) 18.00 18.00 18.00
Exchange Rate: (US$/Cdn$) 0.77 0.77 0.77
Metal Production LOM      
Gold (Moz) 14.3
Copper (Mlb) 14,300
Silver (Moz) 68.2
Molybdenum (Mlb) 13.8
Pre-Tax Results:
Net Cash Flow (US$ Billion) 29.8 19.4 40.9
NPV @ 5% Discount Rate (US$ Billion) 9.7 5.8 13.9
Internal Rate of Return (%) 24.0 17.4 30.4
Payback Period (Years) 4.7 7.5 3.9
Post-Tax Results:
Net Cash Flow (US$ Billion) 18.5 11.9 25.6
NPV @ 5% Discount Rate (US$ Billion) 5.8 3.3 8.4
Internal Rate of Return (%) 18.9 13.5 24.0
Payback Period (Years) 6.2 8.7 4.4

 

The 2022 PEA financials are more sensitive to changes in copper price and exchange rate than changes in capital costs and operating costs.

 

1.15Conclusions

 

The 2022 PFS shows robust economics for KSM ramping up to 195,000 t/d of mill feed from 33 years of open pit mining at the Mitchell, East Mitchell, and Sulphurets deposits to produce a copper-gold-silver concentrate, gold-silver doré, and a molybdenum concentrate.

 

The 2022 PEA shows a stand-alone mine plan for a potential future expansion of the KSM mine to the copper rich Iron Cap and Kerr deposits after the 2022 PFS mine plan has been completed.

 

1.16Recommendations

 

It is recommended that Seabridge focus on advancing development of the KSM Property as described in the 2022 PFS by completing the data collection required to conduct a Feasibility Study. The majority of the US$21.7 million to US$27.3 million of estimated Feasibility Study data collection is related to geotechnical site investigations for TMF, site infrastructure, mine water management tunnels, and water storage dam.

 

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2.0Introduction

 

2.1Overview

 

The KSM Property is currently 100% owned by Seabridge. Description and location of the KSM Property is presented in Section 4.0 of this report.

 

2.2Terms of Reference

 

This report was prepared for Seabridge to summarize the results of the 2022 PFS and the 2022 PEA. The 2022 PEA is a stand-alone, conceptual mine plan that has been undertaken to evaluate a potential future expansion of the KSM mine to the Iron Cap and Kerr deposits after the 2022 PFS mine plan has been completed. None of the Mineral Resources incorporated into the 2022 PEA mine plan have been used in the 2022 PFS mine plan.

 

The report includes an updated Mineral Resource estimate of the Mitchell and East Mitchell deposits, and confirmation that the Mineral Resources for the Kerr, Sulphurets, and Iron Cap deposits remain current. The Qualified Persons (QPs) that authored the report are independent of Seabridge Gold Inc. and the Property.

 

The 2022 PFS and 2022 PEA have been prepared for Seabridge based on work performed by the following independent consultants:

 

·Tetra Tech

 

·Wood

 

·MMTS

 

·KCB

 

·Brazier

 

·ERM

 

·BGC (2022 PFS only)

 

·WSP Golder (2022 PEA only)

 

2.2.1Mineral Resource Estimate Update

 

The Mineral Resource estimate is current as of the effective date of this Report and is presented in Section 14.0 of this report.

 

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2.2.22022 PFS and 2022 PEA Press Releases

 

The results of the 2022 PFS and 2022 PEA were disclosed in Seabridge’s press releases dated June 28, 2022, and August 03, 2022, respectively. This Report is filed in support of the disclosure of the 2022 PFS and 2022 PEA results.

 

2.3Sources of Information

 

The key information sources for the Report were:

 

·documents referenced in Section 3.0 (Reliance on Other Experts)

 

·documents referenced in Section 27.0 (References) of this Report

 

·additional information provided by Seabridge personnel where required.

 

2.4Effective Dates

 

The Report has several effective dates as follows:

 

·KSM Mineral Resource estimate for Mitchell and East Mitchell: March 31, 2022

 

·KSM Mineral Resource estimate for Kerr, Sulphurets and Iron Cap: December 31, 2019

 

·KSM Mineral Reserve estimate: May 26, 2022

 

·the overall effective date of the Report is August 08, 2022.

 

2.5Qualified Persons

 

The name of the QPs of this report and their QP certificates are included in Section 28.0.

 

2.6Personal Inspections

 

The following QPs conducted a site visit of the Property:

 

·Derek Kinakin (M.Sc., P.Geo., P.G.) of BGC visited the Property from August 13 to 17, 2018 and from August 19 to 21, 2013. The site visit involved a general review of the site conditions for the Mitchell and Sulphurets deposits, and detailed inspection and qualitative geotechnical assessment of the core from many of the holes that had been drilled up to that time.

 

·Hassan Ghaffari (P.Eng.) of Tetra Tech visited the Property on September 20, 2014, and conducted a general site overview in the proposed PTMA area.

 

·James H. Gray (P.Eng.) of MMTS visited the Property from June 21 to 22, 2017, and also during 2008, 2009 and 2010. The 2008, 2009 and 2017 visits were general site overview of the potential open pit and rock storage sites and water management structures. Site access alternatives and the route of the MTT were also viewed. The 2010 visit was specifically to inspect the open pit mining areas before the winter snowpack receded.

  

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·Henry Kim (P. Geo) of Wood visited the Property from May 24 to 26, 2022. During the site visit, examined drill cores from all KSM sites, reviewed geology presentation by Seabridge geologists, completed an aerial overview of the Property , and discussed the drill hole database and QA/QC procedures with Seabridge geologists.

 

·Jianhui (John) Huang (Ph.D., P.Eng.) of Tetra Tech visited the Property from June 21 to 22, 2017, from April 27 to 28, 2017 and on September 16, 2008, and conducted an overview of the proposed general PTMA and process plant sites and inspected the drill cores. He also visited ALS Metallurgical laboratory several times during the last 10 years for KSM, including the most recent visit on November 2, 2021 to witness metallurgical testing and samples.

 

·Neil Brazier (P.Eng.) of Brazier visited the Property from September 12 to 16, 2011, from July 20 to 25, 2012, from September 1 to 4, 2013 and from October 4 to 7, 2021 and inspected transmission line routes, the plant site area and locations of energy recovery projects.

 

·David Willms (P.Eng.) of KCB visited the Property on May 24, 2022. The purpose of the site visit was to gain familiarity with the site to support engineering studies and field investigations. Access was limited due to snow cover. Key area of interest that were visited were: Mitchell Valley – including the RSF area and the diversion portals, Sulphurets Valley, the PTMA area, the TMF area, and the WSD area.

 

·Rolf Schmitt (P.Geo.) of ERM visited the Property on July 12, 2019, via helicopter access accompanied by Jessy Chaplin, Seabridge Director, Permitting. The visit entailed a fly-over of the PTMA area, including Treaty Creek access road and powerline route from Bell-Irving River. The site visit also included fly-over of Iron Cap Exploration adit area and ground visit in upper Mitchell Creek valley at the glacier terminus, Camp 9, and the KSM exploration camp at Sulphurets Lake to examine geochemical kinetic field tests and examples of drill core from ICEA adit area.

 

·Ross D. Hammett (Ph.D., P.Eng.) of WSP Golder visited the Property from September 12 to 15, 2017 and from August 8 to 10, 2010 and from October 18 and 19, 2011. The 2017 site visit involved a general inspection of the site conditions, with a specific focus on the stability of the mountain slopes where portals are proposed to be constructed, and all aspects of surface water management related to the proposed mining. It also involved visits to the drill rigs where core was being logged, and detailed inspection and qualitative geotechnical assessment of the core from many of the holes that had been drilled since 2012. The 2010 and 2011 site visits involved a general review of the site conditions for the Mitchell, Iron Cap, and Sulphurets deposits, and detailed inspection and qualitative geotechnical assessment of the core from many of the holes that had been drilled up to that time.

 

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3.0Reliance on Other Experts

 

Ownership, Mineral Tenure, Property Agreements, Surface Rights and Royalties

 

Mr. Henry Kim, P.Geo., Mr. Jim H. Gray, P.Eng., Mr. Ross Hammett, P.Eng., and Mr. Hassan Ghaffari, P.Eng., relied on and disclaims responsibilities for information provided by:

 

·Mr. John Brassard, President of The Claim Group Inc. (TGC), for matters relating to mineral and placer claims status and ownership. The reliance is based on a letter from TGC to Seabridge titled “Seabridge Gold Inc., Title Review-KSM Property, Province of British Columbia”, dated July 21, 2022. Mr. Henry Kim, who is responsible for the information in Section 4, has fully relied on the information provided by Mr. Brassard regarding the claims which comprise the KSM property, their ownership, and their status in Section 4. Mr. Jim H. Gray entirely relied on the information in Section 15.0 and 22.15. Mr. Ross Hammett entirely relied on the information in Section 22.15. Mr. Hassan Ghaffari entirely relied on the information in Section 22 and 24.22.

 

·Seabridge for matters relating to property agreements, royalties, metal streaming agreements, and encumbrances that apply to the KSM Property, Mr. Henry Kim fully relied on the information in Section 4, Mr. James H. Gray entirely relied on the information in Section 15.0 and 22.15. Mr. Ross Hammett entirely relied on the information in Section 22.15. Mr. Hassan Ghaffari entirely relied on the information in Section 22.0 and 24.22. The reliance is based on a letter from Seabridge titled “Seabridge Gold Inc. (“Seabridge”) – Technical Report S. 4.0 Confirmation of Property Matters” dated August 4, 2022 authored by C. Bruce Scott, Vice President General Counsel and Corporate Secretary.

 

Marketing Studies and Contracts

 

Dr. John Huang, Ph.D., P.Eng. relied on:

 

·Mr. Neil Seldon of Neil S. Seldon & Associates Ltd. (NSA) for matters relating to the smelting terms, refining terms, saleability, and sales terms for copper concentrate and molybdenum concentrate Updated in May 2022. These terms are summarized in Section 19.0 and 24.19 and applied in Section 22 and 24.22.

 

Royalties

 

Mr. Hassan Ghaffari, P.Eng. relied on:

 

·Seabridge management concerning private royalties applicable to the PFS which is applied in Section 22.0, and to the PEA which is applied in 24.22.

 

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Taxes

 

·Mr. Hassan Ghaffari has not independently reviewed the taxation information.

 

·Mr. Hassan Ghaffari has fully relied upon, and disclaim responsibility for, taxation information derived from experts retained by Seabridge contained in the following document:

 

̶A letter authored by PricewaterhouseCoopers LLP (“PwC”) with the title:

 

̶“NI 43-101 Technical Report Prepared for Seabridge Gold Inc. – Taxation Narrative” dated August 4, 2022.

 

̶PwC is an Ontario limited liability partnership, which is a member firm of PricewaterhouseCoopers International Limited, each member firm of which is a separate legal entity.

 

̶This information is used in Section 22.4 and 24.22.2 of the Report.

 

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4.0Property Description and Location

 

4.1Mineral Tenure

 

The KSM Property is in northwest BC at an approximate latitude of 56.50 N and a longitude of 130.30 W. The Mineral Resources that are the subject of this report are located relative to the North American Datum (NAD)83 Universal Transverse Mercator (UTM) coordinate system. The Property is approximately 950 km northwest of Vancouver, 65 km north-northwest of Stewart, and 21 km south-southeast of the Eskay Creek Mine (production ceased in 2009). Figure 1.2 is a general location map.

 

The KSM Property comprises five discrete claim blocks and a group of placer claims. Placer claims are on only part of the westernmost claim block of the KSM Property. Claim blocks of the KSM Property are referred to as:

 

·the KSM claims

 

·the Seabee claims

 

·the Tina claims

 

·the Treaty Creek Switching Station claims

 

·the East Mitchell (Snowfield) claim

 

The five KSM claim blocks include 80 mineral claims (cell and legacy) and 2 mining leases with a combined area of 42,052.40 ha. There are also 17 KSM placer claims held by KSM Mining ULC covering part of the KSM claims. The placer claims secure rights in a historically designated placer district. TCG acts as Agent on behalf of Seabridge with respect to maintaining all pertinent records associated with the KSM Property tenures. All claims and leases are in good standing under the Mining Tenure Act of BC and are recorded as owned 100% by KSM Mining ULC, a wholly owned subsidiary of Seabridge.

 

The Seabee and Tina claim blocks are located about 19 km northeast of the Kerr-Sulphurets-Mitchell-Iron Cap mineralized zones. These claim blocks are currently being considered for proposed infrastructure siting. The Treaty Creek Switching Station claims, adjacent to the NTL and east of the Seabee claims are where BC Hydro is constructing the Treaty Creek Terminal (TCT) that will connect KSM to the NTL.

 

The initial group of KSM mineral claims were purchased by Seabridge from Placer Dome in 2001. The mineral claims were converted from legacy claims to BC’s new Mineral Titles Online (MTO) system in 2005. In the MTO system, claims are located digitally using a fixed grid on lines of latitude and longitude with cells measuring 15 seconds north-south and 22.5 seconds east-west (approximately 460 m by 380 m). The legacy claims were located by previous owners by placing tagged posts along the boundaries; however, the survey method employed in locating the legacy claims is not known. The MTO system, where no markings are required on the ground, eliminates the potential for gaps and/or overlapping claims inherent in the old system.

 

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There is no record or evidence of any significant historical mining on the Property other than minimal placer mining. There are no known environmental liabilities to which the Property is subject apart from current exploration activity that is fully permitted. See Section 20 for more details on permits and environmental liabilities. The BC Mineral Inventory (Minfile) contains 25 mineral occurrences associated with previous work in this area (mostly copper and gold).

 

Since it acquired the original KSM claim group, Seabridge has expanded its property holdings around the original KSM claims through staking and purchase of several claim groups. These groups include the Seabee group, acquired by staking, the Tina and BJ groups purchased in 2009, the New BJ group purchased in 2011, the Treaty Creek Switching Station claims purchased in 2018, and the East Mitchell claim purchased in 2020. Additional fractional claims within the Seabee claim block were purchased in 2017. . The Seabee, Tina, and Treaty Creek Switching Station claims are together referred to as the Seabee Property, and the original KSM group, BJ, and New BJ groups are referred to as the KSM Property (Figure 4.1). In 2014, most of the original KSM claims and the BJ claims acquired in 2009 were converted into two mining leases. The 17 KSM placer claims are shown in Figure 4.2.

 

Annual holding costs for all leases and claims vary by year depending on whether the fees are paid in cash or the value of work completed on developing the claims is used in lieu of a cash payment. Over the next five years, the annual cash holding costs to keep the claims and leases valid range between Cdn$250,000 to Cdn$300,000. Those estimated costs can be reduced significantly if work expenditures are applied in lieu of cash fees. No additional permits are required to address the recommendations in this report; part of the expenditures for that work can be applied in lieu of cash fees.

 

Mr. Kim has relied on information with respect to all mining claim matters as provided by TCG in a letter titled “Seabridge Gold Inc. Title Review – KSM Property, Provide of British Columbia”, by John Brassard, dated July 21, 2022.

 

Seabridge’s mineral claim blocks including the KSM, Seabee, Tina, Treaty Creek Switching Station, and East Mitchell groups are shown in Table 4.1 and Table 4.3. The location of the five mineralized zones (Kerr, Sulphurets, Iron Cap, Mitchell, and East Mitchell) is depicted in the southwestern portion of Figure 4.1. Table 4.2 shows Seabridge’s placer claims.

 

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Figure 4.1 KSM Mineral Claim Map

 

 

  

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Figure 4.2KSM Placer Claim Map

 

 

 

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Table 4.1KSM Mineral Claims and Leases

 

Claim
No.
Claim
Name
Area
(ha)
Good To
Date
Map
Number
Mineral Claims (20)
394780 BJ 5 100.0000 25-Nov-2031  104B059
394781 BJ 6 100.0000 25-Nov-2031  104B059
394786 BJ 11 500.0000 25-Nov-2031  104B059
394787 BJ 12 500.0000 25-Nov-2031  104B059
394788 BJ 13 100.0000 25-Nov-2031  104B059
394789 BJ 13A 25.0000 25-Nov-2031  104B059
394790 BJ 14 100.0000 25-Nov-2031  104B059
394791 BJ 15 250.0000 25-Nov-2031  104B059
394794 BJ 18 300.0000 25-Nov-2031  104B059
394808 BJ 31A 375.0000 25-Nov-2031  104B049
394809 BJ 32 150.0000 25-Nov-2031  104B049
394810 BJ 33 450.0000 25-Nov-2031  104B049
394811 BJ 34 150.0000 25-Nov-2031  104B049
394812 BJ 35 450.0000 25-Nov-2031  104B049
683463 - 1,246.5185 25-Nov-2031  104B
683483 - 837.5991 25-Nov-2031  104B
705591 BJ GAP1 231.6166 05-Feb-2032  104B
705592 BJ GAP2 160.4624 05-Feb-2032  104B
1036269 KSM 1  53.4100 21-May-2023  104B
1036270 KSM 2  17.8100 21-May-2023  104B
Totals - 6,097.4166 - -
Mineral Leases (2)
1031440 - 6,085.0000 06-Oct-2023 -
1031441 - 5,162.0000 06-Oct-2023 -
Totals - 11,247.0000 - -
Mineral Claim (1)
509216 - 1267.43 31-Jan-2031  104B
Total - 1267.43    

 

 

Seabridge Gold Inc.4-5221062-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
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Table 4.2KSM Placer Claims

 

Claim
No.
Claim
Name
Area
(ha)
Good to
Date
Map
Number
Placer Claims (17)
986922 PL21 35.7243 16-May-2023 104B
986924 PL22 35.7244 16-May-2023 104B
986925 PL23 107.1670 16-May-2023 104B
541785 Mitchell Placer 3 178.6000 10-Mar-2023 104B
542065 Mitchell Sulphurets Junct. 71.4400 10-Mar-2023 104B
543053 Add-On Placer 2 17.8600 10-Mar-2023 104B
543054 Add-On Placer 3 17.8500 10-Mar-2023 104B
558630 South Sulphurets Placer 71.4600 10-Mar-2023 104B
575058 Mitchell Placer 1 & 2 499.7600 10-Mar-2023 104B
577710 2008 Add-On Placer 17.8600 10-Mar-2023 104B
577712 2008 Add-On Placer 2 17.8600 10-Mar-2023 104B
577715 2008 Add-On Placer 3 17.8600 10-Mar-2023 104B
577716 2008 Add-On Placer 4 17.8600 10-Mar-2023 104B
578100 Sulphuret Creek Placer 2 17.8600 10-Mar-2023 104B
578154 Lower Sulphuret Creek 285.8200 10-Mar-2023 104B
578160 Lower Sulphuret Creek 17.8600 10-Mar-2023 104B
583353 Unnamed & Add-On Placer 1 124.9900 10-Mar-2023 104B
Totals - 1,553.5557 - -

 

Seabridge Gold Inc.4-6221062-01-RPT-001

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Table 4.3Seabee/Tina KSM Claims

 

Claim
No.
Claim
Name
Area
(ha)
Good To
Date
Map
Number
SEABEE Property Mineral Claims (46)
566467 BRIDGE1 445.8258 08-Apr-2028 104A
566468 BRIDGE2 445.5733 08-Apr-2028 104A
566469 BRIDGE3 427.7919 08-Apr-2028 104A
566470 BRIDGE4 427.9770 08-Apr-2028 104A
566471 BRIDGE5 445.7336 08-Apr-2028 104A
566472 BRIDGE6 445.5766 08-Apr-2028 104A
566473 BRIDGE7 427.9217 08-Apr-2028 104A
566474 BRIDGE8 427.7599 08-Apr-2028 104A
566475 BRIDGE9 427.6131 08-Apr-2028 104A
566476 BRIDGE10 445.5312 08-Apr-2028 104A
566477 BRIDGE11 302.8823 08-Apr-2028 104A
566478 BRIDGE12 427.4311 08-Apr-2028 104A
566479 BRIDGE13 445.1533 08-Apr-2028 104A
566481 BRIDGE14 445.0611 08-Apr-2028 104A
566482 BRIDGE15 444.8427 08-Apr-2028 104A
566484 BRIDGE16 444.5621 08-Apr-2028 104A
566485 BRIDGE17 426.7283 08-Apr-2028 104A
566487 BRIDGE18 444.7114 08-Apr-2028 104A
566488 BRIDGE19 444.8346 08-Apr-2028 104A
566489 BRIDGE20 444.9690 08-Apr-2028 104A
566490 BRIDGE21 427.2642 08-Apr-2028 104A
566491 BRIDGE22 445.1671 08-Apr-2028 104A
566492 BRIDGE23 427.3078 08-Apr-2028 104A
566493 BRIDGE24 427.9239 08-Apr-2028 104A
566494 BRIDGE25 427.9246 08-Apr-2028 104A
566495 BRIDGE26 444.8785 08-Apr-2028 104A
566496 BRIDGE27 391.3145 08-Apr-2028 104B
566497 BRIDGE28 444.4573 08-Apr-2028 104A
566567 BRIDGE29 427.4572 08-Apr-2028 104A
571582 SEABEE1 408.8286 08-Apr-2028 104A
571583 SEABEE2 373.1368 08-Apr-2028 104A
571584 SEABEE3 444.0680 08-Apr-2028 104A
571585 SEABEE4 426.0832 08-Apr-2028 104A
571586 SEABEE5 372.6392 08-Apr-2028 104A
571587 SEABEE6 159.6419 08-Apr-2028 104A
573813 SEABEE7 213.2634 08-Apr-2028 104A
575633 SEA 1 445.1987 08-Apr-2028 104A
575635 SEA 2 445.3012 08-Apr-2028 104A
575636 SEA 3 445.4096 08-Apr-2028 104A
575638 SEA 4 445.4484 08-Apr-2028 104A
      table continues…

 

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Claim
No.
Claim
Name
Area
(ha)
Good To
Date
Map
Number
575639 SEA 5 445.3365 08-Apr-2028 104A
575642 SEA 6 445.0850 08-Apr-2028 104A
575643 SEA 7 213.4398 08-Apr-2028 104A
575645 SEA 8 427.0822 08-Apr-2028 104A
575646 SEA 9 35.5980 08-Apr-2028 104A
603133 SEABEE 8 426.5614 08-Apr-2028 104B
Totals - 18,674.2970 - -
 
TINA Property Mineral Claims (7)
401548 TINA 1 500.0000 28-Apr-2032 104B070
401549 TINA 2 500.0000 28-Apr-2032 104B070
401550 TINA 3 500.0000 28-Apr-2032 104B070
401551 TINA 4 500.0000 28-Apr-2032 104B070
401552 TINA 5 500.0000 28-Apr-2032 104B070
401553 TINA 6 250.0000 28-Apr-2032 104B070
603134 SEABEE 9 53.3796 28-Feb-2032 104B
Totals - 2,803.3796 - -
 
Treaty Creek Switching Station Claims (6)
1044111 -  71.1200 14-Nov-2031 104B
1044114 -  106.7800 14-May-2032 104B
1044281 -  53.3700 23-Nov-2031 104B
1047928 -  17.7900 17-Nov-2031 104B
1064020 -  53.4100 20-Nov-2022 104A
1064024 -  106.8400 20-Nov-2022 104A
Totals - 409.3100 - -

 

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4.2Royalties

 

4.2.1KSM Property Royalties

 

The original KSM claim group consisted of two contiguous claim blocks known as the Kerr and Sulphurets (or Sulphside) properties. These claims are 100% owned by the Issuer through its wholly-owned subsidiary, KSM Mining ULC. Newmont Corporation retains a 1% NSR royalty that is capped at Cdn$4.5 million.

 

Two of the pre-converted claims (Xray 2 and Xray 6), the areas of which have now been converted into part of mining lease 1031440, and one pre-converted claim (Xray 8), the area of which is now within the East Mitchell Property (mineral claim 509216) are also subject to an effective 1% NSR royalty capped at US$650,000 and Seabridge has been paying, annually, minimum advance royalty payments that may be credited against future royalties.

 

In addition, the Issuer has granted two options to a subsidiary of Royal Gold, Inc. under which such subsidiary can acquire a 1.25% NSR Royalty and a 0.75% NSR Royalty on gold and silver produced from the KSM Property for Cdn$100 million and

Cdn$60 million, respectively, subject to certain conditions.

 

The Treaty Creek Switching Station Claims and certain fractional claims within the Seabee claims are subject to royalties, however none of the Mineral Resources at the KSM Project are located on the claims subject to these royalties and they are intended for infrastructure siting.

 

Under the Benefits Agreement with the Nisga’a Nation and the Co-operation and Benefits Agreement with the Tahltan Nation, the Issuer has agreed to pay each Nation annual payments. The combined annual payments to these Nations are payable in two forms; payments that are a percentage of the tax payable (the “Mineral Tax”) under the Mineral Tax Act (British Columbia) (the “Mineral Tax Act”), which is a tax on net operating profit of the KSM Project, and payments that are based on net smelter returns of the KSM Project.

 

Payments under the IBA are included in the economic analysis in Section 22 and 24.

 

4.2.2Sprott and Teachers Royalties

 

In March, 2022, KSM Mining ULC (“KSMCo”), a wholly-owned subsidiary of Seabridge, sold a US$225 million secured note (the “Note”) that is to be exchanged at maturity for a silver royalty on the Project and Seabridge sold concurrently a Contingent Right to Sprott Private Resource Streaming and Royalty (B) Corp. (“Sprott”) for US$225 million (approximately Cdn$285 million at the exchange rate at the time). The proceeds of this sale will be used to fund the works KSMCo is planning to advance the Project towards the designation of ’substantially started’.

 

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The principal terms of the Note and silver royalty include:

 

(i)The Note matures on the date (the “Maturity Date”) that is the first to occur of:

 

a.commercial production being achieved at KSM; and

 

b.either March 24, 2032, or, if the environmental assessment certificate (the “EAC”) expires and Sprott does not exercise its right to put the Note to KSMCo (described below), March 24, 2035,

 

and Sprott has agreed it will use all of the principal amount repaid on the Maturity Date (i.e., US$225 million) to purchase a 60% gross silver royalty (the “Silver Royalty”) on the KSM Project.

 

(ii)Prior to the Maturity Date, the Note bears interest at 6.5% per annum, payable quarterly in arrears. KSMCo can elect to satisfy interest payments in cash or by having Seabridge issue common shares of equivalent value under the Contingent Right.

 

(iii)KSMCo has the option to buyback 50% of the Silver Royalty, once purchased by Sprott, on or before 3 years after commercial production has been achieved, for an amount that provides Sprott a minimum guaranteed annualized return.

 

(iv)If project financing to develop, construct and place the KSM Project into commercial production is not in place by March 24, 2027, Sprott can put the Note back to KSMCo for US$232.5 million, with KSMCo able to pay such amount in cash or by having Seabridge issue common shares under the Contingent Right, at KSMCo’s option. This put right expires once such project financing is in place. If Sprott exercises this put right, its right to purchase the Silver Royalty terminates.

 

(v)If KSM’s EAC expires at anytime while the Note is outstanding, Sprott can put the Note back to KSMCo for US$247.5 million at any time over the following nine months, with KSMCo able to satisfy the put in cash or by having Seabridge issue common shares under the Contingent Right, at KSMCo’s option. If Sprott exercises this put right, its right to purchase the Silver Royalty terminates.

 

(vi)If commercial production is not achieved at KSM prior to March 24, 2032, the Silver Royalty payable to Sprott will increase to a 75% gross silver royalty (if the EAC expires during the term of the Note and the corresponding put right is not exercised, this increase in the royalty percentage will occur if commercial production is not achieved at KSM prior to March 24, 2035).

 

(vii)No amount payable may be paid in common shares of Seabridge if, after the payment, Sprott would own more than 9.9% of Seabridge’s outstanding shares.

 

(viii)KSMCo’s obligations under the Note are secured by a charge over all of the assets of KSMCo and a limited recourse guarantee from Seabridge secured by a pledge of the shares of KSMCo.

 

(ix)The KSM project is expected to be operated through joint ownership of KSMCo, and if this occurs the Note and Silver Royalty would thereby become obligations of a future joint venture to develop KSM.

 

Seabridge believes they have addressed all issues to secure access, mineral title, and ability to perform work on the property and are not aware of any risks, other than those identified in this Report, that could materially affect proposed work plans.

 

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5.0Accessibility, Climate, Infrastructure, Local Resources and Physiography

 

The following section was taken from RMI’s April 6, 2007 NI 43-101 report entitled “Mitchell Creek Technical Report, Northern British Columbia” (Lechner, 2007), remains largely unchanged, and has only been updated for consistency in abbreviations and grammar.

 

The Property lies in the rugged Coastal Mountains of northwestern BC, with elevations ranging from 520 m in Sulphurets Creek valley to over 2,300 m at the highest peaks. Valley glaciers fill the upper portions of the larger valleys from just below the tree line and upwards. The glaciers have been retreating for at least the last several decades. Aerial photos indicate the Mitchell Glacier has retreated more than 1 km laterally and perhaps several hundred metres vertically since 1991.

 

The Property is drained by Sulphurets and Mitchell Creek watersheds that empty into the Unuk River, which flows westward to the Pacific Ocean through Alaska. The tree line lies at about 1,240 masl, below which a mature forest of mostly hemlock and balsam fir abruptly develops. Fish are not known to inhabit the Sulphurets and Mitchell watersheds. Large wildlife such as deer, moose, and caribou are rare due to the rugged topography and restricted access; however, bears and mountain goats are common.

 

The climate is generally that of a temperate or northern coastal rainforest, with sub-arctic conditions at high elevations. Precipitation is high with annual rainfall and snowfall totals estimated to be somewhere between the historical averages for the Eskay Creek Mine and Stewart, BC. These range from 801 mm to 1,295 mm of rain and 572 cm to 1,098 cm of snow, respectively (data to 2005). The length of the snow-free season varies from about May through November at lower elevations and from July through September at higher elevations. Exploration activities have typically been carried out from late May into November. It is envisioned that operations would be conducted throughout the year with assets required for snow removal.

 

Currently, access to the Property is via helicopter. Three staging areas for mobilizing crews and equipment were used. These are:

 

1.An area located at kilometre 54 on the private Eskay Creek Mine Road, which is about 25 km to the north-northwest of the Property.

 

2.Along the public Granduc Road, which is located about 35 km to the south-southeast of the Property, which in turn is about 40 km north of the town of Stewart, BC. A section of this road passes through Alaska and the town of Hyder. This area has not been utilized since 2011.

 

3.The Bell 2 Lodge, on Highway 37, 40 km east-northeast.

 

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Construction is currently underway on KSM’s 30-km long TCAR that is expected to reach the KSM PTMA and connect PTMA to Highway 37 in 2023, and the 33-km long CCAR that connects the mine area to Highway 37 via the 59 km long Eskay Creek mine resource road.

 

KSM will connect to BC Hydro’s existing NTL at BC Hydro’s Treaty Creek Switching Station (“TCT”). This TCT, located adjacent to the NTL and Highway 37, 18 km south of Bell 2 Lodge, is scheduled to be completed at the end of 2024. KSM Mining has completed its design for a 30 km long 287 KV transmission line to interconnect the TCT and the KSM plant site. This KSM transmission line is scheduled to be constructed in 2023 with completion and commissioning planned for late 2024 to be ready for connection to the TCT.

 

Stewart, a town of approximately 500 inhabitants, is the closest population center to the Property. It is connected to the provincial highway system via paved, all weather Highway 37A. The larger population centers of Prince Rupert, Terrace, Kitimat, and Smithers, with a total population of about 36,000, are located approximately 270 km to the southeast.

 

There are multiple deep-water loading facilities for shipping bulk mineral concentrates located in the ice-free Port of Stewart, BC. Those port facilities are currently used by the Red Chris Mine. The nearest railway is the CNR Yellowhead route, which is located approximately 220 km southeast of the Property. This line runs east-west and can deliver concentrate to deep-water ports near Prince Rupert and Vancouver, BC.

 

The Property lies on Crown land; therefore, all surface and access rights are granted by the Mineral Tenure Act, the Mining Right of Way Act, and the Mining Rights Amendment Act. There are no settlements or privately owned land in this area; there is limited commercial recreational activity in the form of helicopter skiing and guided fishing adventures. The closest power transmission lines run along the Highway 37, 40 km east of the Property, and along the 37A corridor to Stewart, approximately 50 km southeast.

 

Newcrest (formerly Pretium) is currently mining their high-grade underground Brucejack deposit that is located about 5 km east of the Property. The Brucejack mine site is accessible via a year-round, all weather road branching from Highway 37.

 

Additional details on availability of power, water, waste rock and tailings storage areas and surface rights to support mine operations are discussed in the relevant sections of this Report.

 

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6.0History

 

6.1Exploration History

 

The modern exploration history of the Property began in the 1960s, with brief programs conducted by Newmont Exploration of Canada Ltd. (Newmont), Granduc, Phelps Dodge, and the Meridian Syndicate. All of these programs were focused towards gold exploration. Various explorers were attracted to this area due to the numerous large, prominent pyritic gossans that are exposed in alpine areas. There is evidence that prospectors were active in the area prior to 1935. Several short hole, reconnaissance level drilling programs were undertaken between 1969 and 1991. The Sulphurets Zone was first drilled by Granduc Mines in 1968, Kerr by Brinco Ltd. in 1985, Mitchell Creek by Newhawk Gold Mines Ltd. (Newhawk) in 1991, and Iron Cap by Esso Minerals in 1980.

 

In 1989, Placer Dome (Placer) acquired a 100% interest in the Kerr deposit from Western Canadian Mines; in the following year, they acquired the adjacent Sulphurets Property from Newhawk. The Sulphurets Property also hosts the Mitchell Creek deposit and other mineral occurrences. In 2000, Seabridge acquired a 100% interest from Placer in both the Kerr and Sulphurets properties, subject to capped royalties. In 2020, Seabridge acquired a 100% interest from Pretium Resources in the adjacent East Mitchell (Snowfield) project, subject to capped royalties.

 

There is no recorded mineral production, nor evidence of it, from the Property. Immediately west of the Property, small-scale placer gold mining has occurred intermittently in the Sulphurets and Mitchell creeks.

 

Table 6.1 summarizes the more recent exploration history of the Property.

 

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Table 6.1Exploration Summary of the Kerr Zone

 

Year Activity
1982-1883 “Alpha JV” began prospecting and soil geochem surveys of the Kerr gossan focusing on gold
1984-1985 Brinco optioned the Kerr project, completed some geologic surveys and drilled 3 holes
1987-1989 Western Canadian Mines optioned Kerr and completed 59 drill holes and recognized Cu-Au porphyry
1989 Placer acquires Kerr property
1990-1992 Placer began delineation drilling of Kerr deposit at 50 m centers by drilling 83 holes totalling 16,414 m
1992-1996 Placer estimated resources (pre NI 43-101), met testwork, and scoping studies
1996-2000 Project was dormant
2000 Seabridge acquired a 100% interest in Kerr from Placer Dome
2002 Noranda Inc. acquired an option from Seabridge with the right to earn up to a 65% interest in Kerr
2003-2004 Noranda Inc. undertook various exploration surveys
2006 Seabridge purchased Falconbridge (formerly Noranda) option
2009 Seabridge drilled 7 holes totaling about 1,159 m, conducted metallurgical testing, and permit work
2010 Seabridge drilled 4 holes totaling about 1,453 m, conducted metallurgical testing, and permit work
2011 Seabridge drilled 9 resource definition holes totaling about 2,631 m, continued with preliminary feasibility studies
2012 Seabridge drilled 5 exploration holes totalling 3,731 m
2013 Seabridge drilled 29 resource definition holes totalling 23,844 m, completed induced polarization and down-hole geophysical surveys
2014 Seabridge drilled 16 resource definition holes totalling 15,909 m, completed magneto-telluric and gravity surveys
2015 Seabridge drilled 5 resource definition holes and 3 geotechnical holes totalling 6,629 m
2016 Seabridge drilled 5 resource definition holes totalling 7,119 m
2017 No exploration
2018 One hole drilled totalling 501 m
2019 Completed airborne ZTEM geophysical survey

 

Table 6.2 summarizes the exploration history of the Sulphurets, Mitchell, and Iron Cap zones.

 

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Table 6.2Exploration Summary of the Sulphurets, Mitchell, Iron Cap Zones, and other Exploration Targets

 

Year Activity
1880-1933 Limited placer gold exploration and mining
1935-1959 Placer gold prospecting and staking of mining claims
1959-1960 Newmont and Granduc conducted surveys including airborne mag. Sulphurets and Iron Cap Au zones discovered. D. Ross, S. Bishop and W. Dawson prospected and stake claims in area
1961-1968 Granduc Mines conducted geologic/geochem surveys, drilled 9 holes into Sulphurets zone. Ross-Bishop-Dawson claims were optioned by Phelps Dodge in 1962, Meridian Syndicate in 1965, and Granduc in 1968
1963 R. Kirkham completed a M.Sc. thesis on the geology of Mitchell and Sulphurets areas
1979-1984 Esso Minerals optioned Sulphurets property and completed early stage exploration including drilling 14 holes (2,275 m)
1985-1991 T. Simpson completed a M.Sc. thesis on the geology of the Sulphurets gold zone
1991 Granduc Mines conducted additional exploration surveys targeting molybdenum and drilled 6 holes into Snowfield zone (Bruceside)
1992 Arbee prospect was optioned by Placer from Newhawk
1991-1992 Newhawk commissioned AB geophysical survey over Sulphurets. Newhawk subdivided Sulphurets property into Sulphside and Bruceside. Placer acquired Sulphside (Sulphurets, Mitchell, Iron Cap, and other prospects)
1992 Placer undertook delineation drilling of Sulphurets deposit at 50 meter centers (23 holes)
1993 J. Margolis completed a PhD thesis on the Sulphurets district. Newhawk-Corona drilled 3 holes in the Snowfields and Josephine zones east of Sulphurets
1992-1996 Placer completed geologic modeling, resource estimation (pre NI 43-101), preliminary met testwork, and scoping studies
1999 Silver Standard Resources acquired Newhawk.
1996-2000 Sulphurets project was dormant
2000 Seabridge acquired a 100% interest in the Sulphurets/Mitchell properties from Placer.
2002 Noranda Inc. acquired an option to earn up to 65% from Seabridge
2003-2004 Noranda Inc. undertook various exploration surveys
2005 Falconbridge Ltd. (formerly Noranda) completed 4,092 m of diamond drilling in 16 holes
2006 Seabridge purchased Falconbridge’s option and drilled 29 holes totaling about 9,129 m at the Sulphurets and Mitchell zones
2007 Seabridge purchased Arbee prospect from D. Ross and drilled 37 holes totaling 15,650 m
2008

Seabridge purchased Arbee prospect from D. Ross, drilled 40 holes totaling 17,328 m, started metallurgical testing, obtained new topographic data, and initiated permit related activities

2009 Seabridge drilled 44 holes totalling 11,844 m (resource definition, geotechnical and water monitoring), conducted metallurgical testing, and intensified permit data collection
table continues…

 

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2010

Seabridge drilled 86 holes totaling 26,756 m (resource definition and geotechnical), conducted metallurgical testing, and intensified permit data collection

2011 Seabridge drilled 54 resource definition holes totalling 18,087 m, continued prefeasibility work, and completed magneto-telluring geophysical surveys
2012 Seabridge drilled 40 holes totalling 18,015 m including 15 holes from other nearby targets (e.g. Camp Zone and Icefield)
2013 Seabridge drilled 11 holes totalling 8,857 m and completed induced polarization and down-hole geophysical surveys
2014 Seabridge drilled 13 holes totalling 13,605 m and completed magneto-telluring and gravity geophysical surveys
2015 Seabridge drilled 3 holes totalling 4,395 m and completed airborne magnetic geophysical surveys
2016 Seabridge drilled one resource definition hole totalling 1,038 m
2016 M.Sc. thesis on the Mitchell Deposit completed by G. Febbo
2017 Seabridge drilled 12 resource definition holes totalling 11,077 m
2018 M.Sc. thesis on the Kerr Deposit completed by S. Rossett
2018 Seabridge drilled 42 holes totalling 25,848 m
2019 Seabridge drilled 26 resource definition holes totalling 6,121 m, completed airborne ZTEM over the entire property and 3D induced polarization surveys at Sulphurets
2021 Ph.D. thesis completed on the KSM District Geology by M. Campbell
2021 Seabridge drilled 3,151.4 m in 9 holes in the Mitchell resource area for geological, geochemical, geotechnical, and metallurgical evaluations in support of resource modelling.

 

Table 6.3 summarizes the exploration history of East Mitchell zone.

 

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Table 6.3Exploration Summary of the East Mitchell (Snowfield) Zone

 

Year Activity
1960-1980 Granduc Mines carried out regional reconnaissance prospecting, mapping, and rock sampling over the entire Sulphurets area resulting in the discovery of several porphyry copper- molybdenum and copper-gold occurrences.
1968 Granduc drilled two holes totalling 711.1 m near the central part of East Mitchell, but grades were considered too low.
1980 Esso Minerals optioned the Sulphurets property and conducted detailed geological mapping, trenching, and rock geochemical sampling. The results of this work led to the discovery of the Snowfield zone.
1981-1983 Esso continued exploring the Snowfield zone which appeared to have the potential for a large, low grade gold deposit.
1983 Esso excavated and sampled 24 trenches, totalling 192 m, in the Snowfield zone outlining a 240 m by 120 m area of gold mineralization with an average grade of 0.088 oz/t gold.
1985 Esso terminated their option of the Sulphurets property. Newhawk and Granduc entered into a 60:40 joint venture agreement with Newhawk operating.
1985-1988 Newhawk tested the Snowfield zone with five diamond drill holes totalling 740 m. At the time, the mineralization was interpreted to be a tabular, shallow, southwardly dipping body averaging 70 m thick.
1988 Preliminary metallurgical testing was carried out on the drill core and prospecting continued on the property until 1989.
1989 Newhawk-International joint venture established a property-wide control grid (8 line-km) and conducted a rock sampling program including further rock sampling and trenching on the Snowfield zone.
1991 Two drill holes, totalling 350 m, tested the Snowfield zone with additional rock sampling along its eastern exposed limits. The Newhawk-International joint venture also funded a doctoral thesis on the property by Jake Margolis, which was published in 1993.
1991 Three deep diamond drill holes, totalling 1,164 m, tested the southern extension of the Snowfield zone and another three drill holes, totalling 295 m, tested the nearby Josephine Vein zone.
1999 Silver Standard acquired the Sulphurets claim through the acquisition of all of the shares of Newhawk, including the subject claims.
2006 Silver Standard evaluated the Snowfield zone with 27 diamond drill holes, totalling 6,141 m, and rock sampling to test the lateral and vertical limits of the gold mineralization. An initial resource estimate was prepared.
2007 Silver Standard evaluated the Snowfield zone with 27 diamond drill holes, totalling 6,141 m, and rock sampling to test the lateral and vertical limits of the gold mineralization. An initial resource estimate was prepared.
2008 Silver Standard drilled 6,945 m in 31 holes. A resource estimate was prepared by Minorex Consulting Ltd.
2009-2010 Silver Standard drilled 23,778 m in 42 drill holes, increased the drill density to 100 m centres in the main body of the Inferred resources outlined in 2008, and extended the known mineralization to the northwest and southeast. Two resource estimates were prepared in 2009 and one in 2010 by P&E.
2010 Silver Standard sells Snowfield Property to Pretium Resources.
2012 Seabridge Gold drilled two holes totalling 454.1 m as part of its geotechnical evaluation in the periphery of the open pit model for the Mitchell Zone.
2018 Seabridge Gold drilled one hole totalling 300 m as part of its geotechnical evaluation in the periphery of the open pit model for the Mitchell Zone.
2020 Seabridge Gold purchased the East Mitchell (Snowfield) Property from Pretium Resources.
2021 Seabridge Gold drilled 3,919.9 m in 10 drill holes for geological, geochemical, geotechnical, and metallurgical evaluations in support of resource modelling. Relogging of all historical drill core was completed.

 

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6.2History of Production

 

There is no known production from the Kerr, Sulphurets, Mitchell and East Mitchell, or Iron Cap deposits.

 

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7.0Geological Setting and Mineralization

 

7.1Geological Setting

 

The KSM porphyry Au-Cu-Mo district is in northwestern British Columbia, within the Intermontane belt of the Canadian Cordillera. Encompassing portions of British Columbia, the Yukon, and Alaska, the Canadian Cordillera is a highly fertile metallogenetic domain that features a suite of ore bodies ranging in age from 1.6 billion to less than 20 million years old (Nelson and Colpron, 2007). The district lies within “Stikinia,” a long-lived volcanic island-arc terrane that was accreted onto the Paleozoic basement of the North American continental margin in the Middle Jurassic. Stikinia is the largest of several fault bounded, allochthonous terranes within the Intermontane belt, which lies between the post-accretionary Tertiary intrusions of the Coast belt and continental margin sedimentary prisms of the Foreland (Rocky Mountain) belt. In the KSM area, Stikinia is dominated by variably deformed, oceanic island arc complexes of the Triassic Stuhini and Jurassic Hazelton groups (Alldrick and Britton, 1988; Kirkham and Margolis, 1995). Back-arc basins formed eastward of the KSM Property in the Late Jurassic and Cretaceous were filled with thick accumulations of fine black clastic sediments of the Bowser Group. Folding and thrusting due to sinistral transpression in the mid-Cretaceous, related to the development of the Skeena fold and thrust belt, followed by extensional conditions generated the area’s current structural features (Febbo et al., 2015). The mid-Cretaceous deformation generated two of the most important structures in the district: the north-northeast striking, moderately west-northwest dipping Sulphurets thrust fault (STF) and MTF. Remnants of Quaternary basaltic volcanic eruptions occur throughout the region.

 

Early Jurassic sub-volcanic intrusive complexes are common in the Stikinia terrane, and several host well-known precious- and base-metal-rich hydrothermal systems. These include copper-gold porphyry zones such as Galore Creek, Red Chris, Kemess, Mt. Milligan, and KSM. In addition, there are several related polymetallic zones including skarns at Premier, epithermal veins, subaqueous vein, and replacement sulphide zones at Eskay Creek, Snip, Brucejack, and Granduc. At KSM, Triassic rocks include marine sedimentary and intermediate volcanic rocks of the Stuhini Group. The lowermost Stuhini Group is dominated by turbiditic argillite and sandstone, which are overlain by volcanic pillowed flows and breccias. The upper portion consists of turbidites and graded sandstones similar to the base strata. The Stuhini Group is separated by an erosional unconformity from the overlying Jurassic sedimentary and volcanic rocks of the Jack Formation and Hazelton Group. The Jack Formation is comprised of fossiliferous, limey mudstones, and sandstones. The base is marked by a granodiorite and limestone cobble bearing conglomerate. Overlying the Jack Formation is the Hazelton Group, dominated by andesitic flows and breccias in a volcanic chain with high paleotopographic relief. Distinct felsic welded tuff horizons of the Mount Dilworth Formation are an important stratigraphic marker in the Hazelton Group, as they are closely associated with the Eskay Creek Zone.

 

A variety of dykes, sills, and plugs of diorite, monzodiorite, granodiorite, monzonite, syenite, and granite are found in the area. U-Pb zircon radiometric dating indicates that most of the intrusions in the district are of Early Jurassic age (~190 – 197 ± 2 Ma; Kirkham and Margolis, 1995; MacDonald, 1993; Bridge, 1993; Febbo et al., 2019a), coinciding with the age of ore formation. Small volumes of thin post-mineral dykes, including mid-Cretaceous diabase dykes and Eocene lamprophyre dykes, are also observed within the Property. The Early Jurassic intrusions in the district are categorized into two regional suites: the calc-alkaline Texas Creek suite and the alkaline Premier suite. Texas Creek suite are intimately related to porphyry Au-Cu-Mo ore formation at the Kerr, Sulphurets, Mitchell and Iron Cap deposits, and are locally referred to as the “Sulphurets suite intrusions.” The Premier suite intrusions are most abundant in the hanging wall of the STF in the district, and comprise sills and plugs of typically coarse-grained porphyritic monzonite to syenite. Figure 7.1 is a generalized geologic map of the KSM district showing lithology, major structures, and mineralized zones. Drill hole locations are shown for all KSM deposits in Section 10.0.

 

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Figure 7.1Geology of the KSM District (Seabridge, 2022)

 

 

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7.2Mineralization

 

7.2.1Kerr Zone

 

The Kerr Zone is centered on an Early Jurassic, north-south trending, steep westerly dipping, tabular intrusive complex. Drilling demonstrates that the Kerr Zone has an extent of 2,400 m along strike, a width of roughly 800 m, and a vertical extent of at least 2,200 m. The flattened and elongated modern morphology of Kerr is highly unusual for porphyry copper-gold deposits, which typically display roughly cylindrical morphologies. The flattened morphology and relatively elevated copper:gold ratio of the Kerr Zone distinguishes Kerr from the other deposits in the district. The surface expression of the Kerr deposit is a large and elongated, northerly trending, pyritic gossan, primarily exposed in a cirque on the northern flank of Kerr peak.

 

The Kerr intrusive complex is composed of a suite of northerly-striking and steeply west-dipping dykes and intrusions emplaced into a sequence of rhythmically bedded siltstones, sandstones, conglomerates, and debris flows belonging to the Lower Jurassic Hazelton Group. Wall rocks adjacent to the intrusions have been hornfelsed and hydrothermally altered but generally contain marginal metal grades. The complex is composed of an east and west limb separated by a thin wedge of intensely altered sedimentary wall rock. The west limb is up to 500 m thick, and the east limb is up to 300 m thick.

 

Mineralization at Kerr was associated with the sequential emplacement of numerous syn-mineral plagioclase-hornblende-(biotite-K-feldspar-apatite)-phyric diorite to monzodiorite dykes, which display a porphyritic texture with 30-40 vol.% phenocrysts up to 5 mm in length. Dykes are typically several meters to several tens of meters wide. Three distinct phases of these intrusions are recognized based upon typical metal content, quartz vein abundance, degree of hydrothermal alteration, and presence of xenoliths of earlier phases. The earliest syn-mineral phase (“P1”) is characterized by strong gold and copper mineralization and a high abundance (50-90 vol.%) of quartz veins. The second phase of syn-mineral diorite intrusions (“P2”) make up the bulk of the Kerr intrusive complex and feature variable mineralization and quartz vein abundance. The P2 intrusions envelop and truncate portions of the P1 intrusions, and are observed to contain occasional P1 xenoliths. Finally, late syn-mineral to post-mineral dykes (“P3”), with weak to no mineralization, cut the earlier phases. Volumetrically minor post-mineral dykes, with true thicknesses typically on the order of a meter, are also observed at the Kerr deposit. These include a set of unmineralized, K-feldspar megacrystic, plagioclase-hornblende-phyric porphyritic dykes, aphanitic diabase dykes, and amygdaloidal lamprophyre dykes. Current geochronological data for the Kerr deposit includes a U-Pb zircon age of 197 ± 3 Ma for a syn-mineral “syenodiorite” intrusion, as well as a slightly younger U-Pb zircon age of 195 ± 1.5 Ma for a late syn-mineral K-feldspar megacrystic porphyry dyke (Bridge, 1993).

 

Small volumes of syn-mineral hydrothermal breccias are also found at Kerr, with subangular to angular clasts of local intrusions or wall rock and rock flour matrices. Hydrothermal breccias with abundant anhydrite and quartz vein fragments (“QABX”) are also observed; these QABX zones are commonly overprinted with strong sericitic alteration and a high-sulphidation bornite – pyrite ± enargite mineral assemblage, and can contain strong copper mineralization. The QABX zone at Mitchell is analogous to the QABX zones at Kerr.

 

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Hydrothermal alteration at the Kerr deposit is consistent with that described for many calc-alkaline porphyry deposits, with zones of propylitic, potassic, sericite-chlorite, and sericitic alteration, as well as small zones of remnant advanced argillic alteration.

 

Post-mineral deformation and faulting, mainly related to the mid-Cretaceous development of the Skeena fold and thrust belt, has significantly impacted the Kerr deposit. Efforts to thoroughly document and reconstruct the geology of the Kerr deposit are hindered by intense hydrothermal alteration and extensive post-mineralization deformation. Large portions of Kerr are affected by texturally destructive sericitic alteration that contains a pervasive foliation, rendering protolith identification difficult, and the occurrence of numerous post-mineral faults further obscure the original spatial relationships within the deposit.

 

Pervasive foliation and vein deformation resulting from mid-Cretaceous deformation is observed throughout the Kerr deposit. However, as noted at the Mitchell and East Mitchell Zones (e.g., Febbo et al., 2015), foliation is strongest in rocks characterized by alteration assemblages with low rock strengths, such as zones with strong sericitic alteration. East-west shortening related to mid-Cretaceous deformation likely contributed to the unusually flattened morphology of the Kerr deposit. Furthermore, the strongest sericitic alteration occurs at the northern end of Kerr, at a lower elevation than the south side of Kerr. Sericitic alteration is typically formed at relatively shallow depths in porphyry copper systems, atop deep potassic alteration. The current emplacement of alteration assemblages suggests tilting or complex structural modification of the system.

 

The dominant copper mineral is chalcopyrite, which typically occurs as isolated grains about 0.2 mm to 2 mm across, disseminated and clustered in quartz veins, fractures, and surrounding haloes. Bornite is present almost exclusively in the north half of the east leg, within a QABX zone containing >50% crackled quartz veins, and is accompanied by coarse grained chalcopyrite and minor tennantite. Tennantite-tetrahedrite is rare, but widely distributed in late quartz-carbonate veins, mostly in wall rocks, along with minor sphalerite, rare galena, and arsenopyrite. Dark, arseniferous pyrite is associated with these minerals. Molybdenite is a very minor constituent, and Kerr contains significantly lower overall Mo grades than the other deposits in the district. Visible gold has not been observed except under microscopic examinations, where it is observed as less than 100 µm inclusions within sulphides, mainly chalcopyrite, and sulphide grain boundaries. High-grade mineralization at Kerr is primarily associated with early potassic alteration and “A-type” and “B-type” quartz veins (using the terminology of Gustafson and Hunt, 1975). At shallower levels in the system, both copper and gold grades are highly correlated with the density of A- and B-type veining. At depth, however, mineralization does not correlate as strongly with quartz vein density, and is of a more disseminated style.

 

A lithological map of the Kerr deposit is shown in Figure 7.2. Representative cross sections through the south and the north of the Kerr deposit, showing lithological units, quartz vein abundance, gold grade, and copper grade, are shown in Figure 7.3 and Figure 7.4, respectively.

 

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Figure 7.2Geological Map of the Kerr Deposit (Seabridge, 2019). Thin dykes are not displayed.

 

 

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Figure 7.3Four panels of the same vertical E-W section through the Kerr deposit, at 6,258,650N, showing different properties of the deposit. (Seabridge, 2019)

 

 

Notes: Location of section is shown on Figure 7.3: a) Simplified lithology, with important intrusive phases (P1, P2, P3, P4, P6), hydrothermal breccia with anhydrite and quartz vein clasts (QABX), hydrothermal breccias alongside P2 intrusions (IBX), and undifferentiated wall rock (WR) shown. Drill hole traces, within ±100 m of the section, are displayed as black lines. Units are described in the text. The intrusion labels refer to early intrusions with high quartz vein densities (50-90 vol.%; P1), syn-mineral plagioclase-hornblende-phyric intrusions (P2), post-mineral plagioclase-hornblende-phyric dykes (P3), post-mineral K-feldspar megacrystic dykes (P4), and post-mineral lamprophyre dykes (P6). b) Isolines of total volume percent quartz vein abundance as logged in drill core. c) Gold grade, from the Seabridge Gold EOY 2016 block model of Kerr. d) Copper grade, from the EOY 2016 block model of Kerr.

 

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Figure 7.4Four panels of the same vertical E-W section through the Kerr deposit, at 6,259,650N, showing different properties of the deposit. (Seabridge, 2019)

 

 

Notes: Location of section is shown on Figure 7.4: a) Simplified lithology, with important intrusive phases (P1, P2, P3, P4, P6), hydrothermal breccia with anhydrite and quartz vein clasts (QABX), and undifferentiated wall rock (WR) shown. Drill hole traces, within ±100 m of the section, are displayed as black lines. Units are described in the text. The intrusion labels refer to early intrusions with high quartz vein densities (50-90 vol.%; P1), syn-mineral plagioclase-hornblende-phyric intrusions (P2), post-mineral plagioclase-hornblende-phyric dykes (P3), post-mineral K-feldspar megacrystic dykes (P4), and post-mineral lamprophyre dykes (P6). b) Isolines of total volume percent quartz vein abundance as logged in drill core. c) Gold grade, from the Seabridge Gold EOY 2016 block model of Kerr. d) Copper grade, from the EOY 2016 block model of Kerr.

 

7.2.2Sulphurets Zone

 

The Sulphurets Zone is situated between the Kerr and Mitchell deposits, immediately to the north of the valley hosting Sulphurets Lake. Sulphurets was historically subdivided into several adjacent mineralized zones, including the Raewyn Copper-Gold, Breccia Gold, Main Copper, and Canyon zones. The main body of the Sulphurets deposit has a lensoidal geometry, dipping approximately 30 degrees northwest. It has a horizontal extent of 2,200 m, down dip extent of 550 m, and true thickness of up to 330 m. Copper and gold grades gradually diminish towards the limits of the known resource area, but the extent of anomalous copper and gold grades has not been completely delineated down dip or in the southwest direction.

 

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Sulphurets is a structurally complex deposit, intersected by numerous east-vergent faults associated with the mid-Cretaceous Skeena fold and thrust belt. The deposit is composed of stacked thrust fault panels made up of Upper Triassic Stuhini Group and Lower Jurassic Hazelton Group volcano-sedimentary strata intruded by several dykes and stocks. The thrust faults and most dykes and intrusions within the Sulphurets Zone strike southwest-northeast and dip at shallow to moderate angles towards the northwest. Three main fault panels make up the deposit, each featuring a distinct set of intrusions and style of mineralization.

 

The uppermost fault panel (“Upper Panel”) is in the hanging wall of the STF. The most voluminous plutonic rocks in the Upper Panel are Premier suite monzonites and syenites, including a large intrusive body, dipping at roughly 40° – 50° degrees towards the northwest, as well as several thinner dykes. The Premier suite intrusions are typically pinkish to reddish K-feldspar-plagioclase-phyric porphyries with abundant coarse, tabular K-feldspar phenocrysts, commonly displaying concentric zoning. A U-Pb zircon radiometric age of 191.8 +6.5/-1.0 Ma was obtained for a Premier “feldspar porphyry” from the Upper Panel (MacDonald, 1993). Most of the Premier intrusions in the Upper Panel display potassic alteration and reddish hematite dusting but are nearly barren of sulphides. Hydrothermal breccias and intensely altered wall rocks peripheral to the intrusions host the bulk of mineralization in the Upper Panel. First discovered in the 1930s, this mineralized zone was originally referred to as the Main Copper Zone. Alteration in the breccias and wall rock ranges from a dark purplish-brown proximal potassic assemblage with hydrothermal K-feldspar, fine-grained biotite, magnetite, and a relatively high chalcopyrite:pyrite ratio, grading outward to a chlorite – pyrite-dominant alteration, to a distal chlorite – epidote – calcite ± magnetite ± hematite propylitic assemblage.

 

The Upper Panel structurally overlies the thin “Middle Panel,” which is bounded by the STF and the Raewyn fault (RF), a minor thrust fault that terminates against the STF on the northeastern side of the deposit. The Middle Panel features Texas Creek (“Sulphurets”) suite plagioclase-hornblende-phyric diorite intrusions and small volumes of hydrothermal breccias, and hosts very little of the mineralization at Sulphurets. Hydrothermal alteration assemblages are limited to chlorite-dominant alteration in the diorite intrusions, and to weak hornfels ± chlorite in the sedimentary wall rocks and hydrothermal breccias along the margins of diorite intrusions.

 

Finally, the “Lower Panel,” which is beneath the Middle Panel, hosts the bulk of the mineralization at Sulphurets. The Lower Panel is thought to be bounded at depth by the MTF, though the MTF is only intercepted in a handful of deep drill holes. This panel was historically divided into several discrete mineralized zones, including, from southwest to northeast: the Canyon zone, the Breccia Gold zone, and the Raewyn Copper-Gold zone (Fowler and Wells, 1995). Much of the gold and copper mineralization in the Lower Panel is found in the Breccia Gold and Raewyn Copper-Gold zones. The Raewyn Copper-Gold zone is primarily hosted in crackled and hornfelsed sedimentary wall rock partially affected by distinctive dark brown, biotite-rich potassic alteration. It has an apparent north-northeast strike and dips about 45° to the northwest, with approximate true dimensions of 1,000 m in strike, 550 m down dip, and up to 250 m in thickness. The Raewyn Copper-Gold zone remains open down dip and along strike to the northeast at depth. The Breccia Gold zone has a pipe-like geometry, approximately 100 m in diameter, and plunging ~50° to the west-northwest. It is thought to be hydrothermal in original, and features tourmaline-rich alteration, strong gold grades, and abundant pyrite and tennantite, but relatively low copper grades. Sedimentary wall rock from the Lower Panel features silica addition as well as potassic alteration, as evidenced through whole rock geochemistry results comparing wall rock from the Lower Panel and the less altered sedimentary wall rock of the Middle Panel.

 

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The Lower Panel features two main types of Texas Creek suite intrusions: subporphyritic plagioclase-hornblende-phyric diorite intrusions, and porphyritic plagioclase-hornblende-K-feldspar-phyric monzodiorite dykes. The diorites typically feature 30-40 vol.% tabular plagioclase and hornblende phenocrysts, up to ~4 mm in length, in a fine-grained groundmass. Similar diorite intrusions are found at the Kerr, Mitchell, East Mitchell, and Iron Cap deposits, where they are intimately associated with porphyry mineralization. However, the diorites at Sulphurets lack abundant porphyry style quartz vein stockworks, have lower pyrite contents, potassic and sericitic alteration facies are subordinate to chlorite-pyrite alteration, and their associated Cu-Au resources are much smaller. A significant proportion of the mineralization at Mitchell, Kerr, and Iron Cap is hosted within syn-mineral diorite intrusions, whereas at Sulphurets, mineralization is in the wall rocks immediately adjacent to, above and between the fingers of diorite.

 

Late, discontinuous veins, breccias, shear fillings, and patchy replacements similar to those in the Upper Panel occur throughout the Lower Panel fault block. A higher density of these late mineralized structures is reflected in average arsenic, lead, antimony, and zinc concentrations being roughly two to three times those of the Upper Panel. Scattered, sub-meter to centimeter scale quartz and sulphide veins, breccia veins, shear fillings, patchy discontinuous clots or replacements and disseminations occur in all rock types except late dykes but are most abundant in brecciated and crackled zones. They include coarse pyrite, carbonate, chlorite, quartz, occasionally with minor chalcopyrite, sphalerite, arsenopyrite or galena, and traces of tennantite. Native gold is rare but has been observed as sub-millimeter blebs and stringers in quartz-carbonate-sulphide veins, and chlorite-pyrite-carbonate breccia matrix fillings.

 

A small number of very thin, volumetrically insignificant, post-mineral dykes with meter-scale widths cut all other lithologies at Sulphurets. These include northwesterly-dipping, fine-grained diabase dykes with aphanitic chilled margins and pervasive chlorite alteration, and two fine-grained, unaltered, black lamprophyre dykes dipping west-northwest. Similar post-mineral dyke sets are observed in the Kerr, Mitchell, and Iron Cap zones.

 

Brittle fracturing is widespread at Sulphurets, and multiple fault sets, ranging from shallow- to steeply-dipping, are found throughout the deposit. In addition to the east-vergent thrust faults, numerous small, north to north-easterly striking, steeply westerly dipping faults, fracture zones and shears cut through the district. Subvertical penetrative cleavage is found throughout much of the district, and is especially well developed in alteration facies rich in phyllosilicates and clays (Kirkham, 1963; Margolis, 1993). At Sulphurets, the absence of significant rheologically weak sericitic alteration resulted in the absence of the strong penetrative cleavage present at the Kerr, Mitchell, and East Mitchell deposits. At Sulphurets, a north-easterly-striking, steeply dipping foliation is only observed in discrete areas with sericitic alteration (e.g., Fowler and Wells, 1995).

 

A geological map of the Sulphurets deposit, which also displays the extent of near-surface gold and copper mineralized zones, is found in Figure 7.5. Representative cross sections through the southwest and the centre of the Sulphurets deposit, showing lithological units, alteration assemblages, gold grade, copper grade, molybdenum grade, and silver grade are shown in Figure 7.6 and Figure 7.7, respectively.

 

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Figure 7.5Map of the Sulphurets deposit (Seabridge, 2019)

 

 

Notes: a) Geology map, showing principal intrusions, dykes, and hydrothermal breccias. Labeled faults include the Sulphurets Thrust Fault (STF), Raewyn Fault (RF), and the approximate trace of the Mitchell Thrust Fault (MTF). b) Map of the principal gold and copper mineralized zones within the Sulphurets deposit, along with the historical names of these zones. The locations of the major fault panels discussed in the text – the Upper, Middle, and Lower Panels – are also indicated.

 

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Figure 7.6Vertical cross-section through the Sulphurets deposit, looking ENE (Seabridge, 2019)

 

Notes: The traces of the Sulphurets Thrust Fault (STF), Raewyn Fault (RF), and Mitchell Thrust Fault (MTF) are shown. a) Simplified geology of the Sulphurets deposit, showing undifferentiated sedimentary wall rock (WR) and principal intrusions and breccias: diorite (DR), monzodiorite intrusions (MZD), hydrothermal breccia (BX), and diabase dyke (DB). See text for unit descriptions. The traces of drill holes within ± 100 m of the section are shown as black lines. b) Simplified diagram of hydrothermal alteration zoning within the Sulphurets deposit, illustrating the extent of pervasive chlorite (CHL), propylitic (PRO), hornfels (HF), and hornfels with tourmaline (TO) alteration zones, as well as areas with quartz-sericite-pyrite overprint (SER). c) Ag grades from the Seabridge Gold EOY 2018 block model; d) Au grades from the EOY 2018 block model; e) Cu grades from the EOY 2018 block model; f) Mo grades from the EOY 2018 block model.

 

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Figure 7.7Vertical cross-section through the Sulphurets deposit, looking NNE (Seabridge, 2019)

 

  

Notes: The traces of the Sulphurets Thrust Fault (STF), Raewyn Fault (RF), and Mitchell Thrust Fault (MTF) are shown. a) Simplified geology of the Sulphurets deposit, showing undifferentiated sedimentary wall rock (WR) and principal intrusions and breccias: diorite (DR), monzonite (MZ), hydrothermal breccia of Middle Panel (BX1), Sulphurets Gold Breccia (BX2), brecciated margin of monzonite intrusions (BX3), and lamprophyre dyke (LM). See text for unit descriptions. The traces of drill holes within ±100 m of the section are shown as black lines. b) Simplified diagram of hydrothermal alteration zoning within the Sulphurets deposit, illustrating the extent of pervasive chlorite (CHL), propylitic (PRO), hornfels (HF), hornfels with tourmaline (TO), hornfels with biotite (BI), potassic (POT; hydrothermal K-feldspar ± biotite ± magnetite), and mixed potassic (POTM) alteration zones. c) Ag grades from the Seabridge Gold EOY 2018 block model; d) Au grades from the EOY 2018 block model; e) Cu grades from the EOY 2018 block model; f) Mo grades from the EOY 2018 block model.

 

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7.2.3Mitchell Zone

 

The Mitchell Zone represents the footwall portion of an Early Jurassic gold-copper-molybdenum porphyry deposit that was cleaved in two by the MTF during the mid-Cretaceous. The upper portion of the deposit, termed the East Mitchell Zone (see Section 7.2.4), was transposed approximately 1,600 m towards the east-southeast during the development of the Skeena fold and thrust belt.

 

The Mitchell Zone crops out in Mitchell Valley, through an erosional window exposing the footwall of the MTF. The zone is a roughly cylindrical gold-copper-molybdenum deposit, dipping ~60° to the northwest, with approximate true dimensions of 1,600 m in strike, 1,500 m down dip, and up to 850 m in thickness. It remains open down dip and along strike to the northeast at depth.

 

Recent glacial melt back has provided exceptional surface exposure of a gold-copper-molybdenum porphyry system. A zone of intense quartz and sulphide veining (“P1” or “Sheeted Quartz Vein Zone”) forms resistant bluffs in Mitchell Valley. However, the higher-grade core area is mostly covered by talus and moraine west of the bluffs. Active oxidation and leaching of sulphides have produced prominent gossans and extensive copper sulphate precipitates at the surface.

 

Mineralization at the Mitchell deposit is genetically and spatially linked to the Early Jurassic Mitchell intrusive complex, which is composed of Sulphurets (Texas Creek) suite diorite, monzodiorite, and granodiorite stocks and dykes. The intrusive complex cuts sedimentary and volcanic rocks of the Upper Triassic Stuhini Group and sandstones, conglomerates, and andesitic rocks of the Lower Jurassic Jack Formation (basal Hazelton Group). The Mitchell complex has been subdivided into three major intrusive phases. “Phase 1” includes voluminous pre- to early-mineral plagioclase-hornblende diorite porphyry intrusions (196 ± 2.9 Ma and 189 ± 2.8 Ma; U-Pb zircon, Febbo et al., 2019) and a suite of westerly-striking and northerly-dipping monzodiorite dykes. “Phase 2” includes a syn-mineral granodiorite plug within the centre of the deposit (192 ± 2.8 Ma; U-Pb zircon, Febbo et al., 2019), which features abundant xenoliths of quartz veins and Phase 1 intrusions, as well as multiple westerly-striking and northerly-dipping granodiorite dykes in the centre and western half of the complex. Following Phase 2 magmatism, a suite of late-mineral breccia dykes was emplaced. Finally, “Phase 3” is made up of a small diorite plug, measuring 50 m by 125 m, which cuts the breccia dykes.

 

The successive intrusive phases were accompanied by the development of different hydrothermal alteration assemblages, veining and mineralization within the Mitchell deposit. Hosted by Phase 1 plutons, Stage 1 sheeted quartz – chalcopyrite – pyrite ± magnetite ± K-feldspar ± chlorite veins and stockworks contain most of the copper-gold mineralization, and are accompanied by proximal potassic to calc-potassic alteration and peripheral propylitic alteration. Stage 2 quartz – pyrite – chalcopyrite – molybdenite veining produced a molybdenum-enriched halo around the centre of the deposit (190.3 ±0.8 Ma; Re-Os molybdenite). The molybdenum halo is associated with sericite-pyrite alteration and is temporally related to a Phase 2 stock that crops out central to the halo. Stage 3 consists of anhydrite veining, poorly mineralized massive pyrite veins, and shallow quartz – pyrophyllite advanced argillic alteration. Chalcopyrite is the predominate copper-bearing mineral, but a discrete bornite-bearing zone is identified just east of the center of the Mitchell deposit. This “Quartz Anhydrite Breccia (QABX)” or “Bornite Breccia” was only intersected in three holes (including one interval of 86 m with 1.42% copper and 0.23 g/t gold), and the interpreted dimensions are about 400 m long down dip, 60 m thick, and 250 m along strike. Its geometry roughly aligns with the northwest plunging trend of the Mitchell deposit. The breccia is composed of a chaotic, swirly mix of crackled and milled light grey quartz, anhydrite, and clay, with disseminated and interstitial pyrite, chalcopyrite, bornite, minor tennantite and molybdenite. In deeper intersects the breccia transitions to a structure containing quartz, anhydrite, pyrite, and chalcopyrite, with only traces of bornite. The breccia body is interpreted to be related to structurally controlled, late advanced argillic alteration. A similar breccia zone, with both bornite and abundant anhydrite, is found in the Kerr deposit.

 

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The Mitchell deposit features many characteristics typical of gold-enriched calc-alkaline porphyry copper deposits, including: calc-alkaline syn-mineral intrusions; hydrothermal alteration assemblages that include deep potassic, peripheral propylitic, and extensive shallow sericitic alteration; and abundant quartz veining. Metals, chiefly gold and copper (in terms of economic value), are generally at low concentrations, finely disseminated, stockwork or sheeted veinlet controlled, and pervasively dispersed over dimensions of hundreds of metres. Grades diminish slowly over large distances; sub-economic grades are encountered at distances of several hundreds of metres beyond the interpreted centre of the system. This is distinct from the Sulphurets and Kerr zones, where there are more abrupt breaks in grade due to higher structural complexity and juxtaposition of weak and moderate grade domains by faulting, both syn-mineral structures controlling breccia contacts, and post-mineral faulting and displacements.

 

Directly above the Mitchell Resource and hanging wall to the MTF are minor occurrences of disseminated and veinlet chalcopyrite in hydrothermal breccias, intrusion breccias, magnetite skarn and hornfels altered sedimentary and volcanic rocks. This limited mineralization is adjacent to non-mineralized Premier suite porphyritic monzonite to syenite intrusions with an age of 193.9 ± 0.5 Ma (Kirkham and Margolis, 1995). This portion of the deposit is more typical of alkalic porphyry deposits, with sodic-potassic alteration, relatively low quartz veining, and K-feldspar – magnetite – albite – pyrite – chalcopyrite – specularite veins. This style of mineralization is similar to that observed in the hanging wall of the STF at the Sulphurets deposit (Main Copper Zone).

 

The deposit was deformed during development of the mid-Cretaceous Skeena fold and thrust belt (SFTB), transforming portions of the deposit into intensely foliated, mylonitic zones, with deformed and flattened quartz veinlets. The main foliation generally strikes to the east-southeast, and dips moderately to the north. The intensity of the foliation and shortening varies within the boundaries of the Mitchell Zone, and is directly related to hydrothermal alteration assemblages. Sections of the deposit featuring rheologically weak rocks, such as areas with intense sericitic alteration, exhibit the most shortening and deformation (Febbo et al., 2015). The MTF, separating the

 

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Mitchell Zone in the footwall from the East Mitchell Zone, also formed during SFTB development (see Section 7.2.4 for further description of the East Mitchell Zone). Deep drill holes have also indicated the presence of a roughly 50 m thick, banded, mylonitic shear zone that may offset the base of the Mitchell deposit, termed the Basal Shear Zone (BSZ). The BSZ dips to the northwest and appears to parallel the MTF. As the Bornite Breccia and BSZ may have structurally offset portions of the Mitchell Zone, potential remains for additional mineralization to be discovered.

 

A geological map of the deposit is presented in Figure 7.8. A representative cross section and 950 masl level plan of the Mitchell and East Mitchell Zones, showing lithological units, alteration assemblages, quartz vein abundance contours, gold grade, copper grade, and molybdenum grade, are shown in Figure 7.9 and Figure 7.10, respectively.

 

Figure 7.8Geology Map of the Mitchell and East Mitchell Deposits (Seabridge, 2022)

 

 

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Figure 7.9Vertical section through the Mitchell and East Mitchell Zones (Seabridge, 2022)

 

 

Notes: Location is shown as A-A’ in Figure 7.10a. The traces of the Mitchell thrust fault (MTF), Mitchell basal shear zone (BSZ), Brucejack fault (BJF), and Snowfield landslide (LAN) surface are shown on each panel. a) Simplified geology of the Mitchell and East Mitchell Zones, showing Stuhini Group undifferentiated sedimentary wall rock (SEDS), Hazelton Group volcano-sedimentary wall rock (VOLC), plagioclase-hornblende diorite porphyry (P2), the sheeted vein body with >50 vol.% quartz veinlets (P1; H1 hybrid zones hosted in wall rock), zones with 25-50 vol.% quartz veinlets (P1B; H1B where hosted in wall rock), syn-mineral plagioclase-hornblende-K-feldspar-phyric porphyry dykes (P3), late syn-mineral porphyry dykes with coarse K-feldspar (K3), monzonite to syenite porphyry (P8), intrusion breccias (P5), post-mineral dolerite dykes (K4), and overburden (OVBD). The traces of drill holes within ± 50 m of the section are shown as black lines. b) Distribution of hydrothermal alteration assemblages within the Mitchell deposit, illustrating the extent of propylitic alteration (PRO), potassic alteration (POT1, POT2), hornfels (HF), hornfels with biotite (HFB), hornfels with tourmaline (HFT), sericitic alteration (QSP), advanced argillic alteration (ADA), as well as the occurrence of hydrothermal anhydrite (ANHY). c) Contours of logged volume percent of total quartz veins within Mitchell drill core. d) Block model Au grades, showing clear concentric zoning; e) block model Cu grades, showing a concentric zoning pattern like that of Au; and f) block model Mo grades, showing a zoning pattern with highest Mo grades rimming the central zone of high Au and Cu in the Mitchell deposit.

 

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Figure 7.10Plan view of the Mitchell deposit at 950 masl (Seabridge, 2022)

 

 

Notes:The traces of the Mitchell thrust fault (MTF), Brucejack fault (BJF), and Snowfield landslide (LAN) surface are shown on each panel. a) Simplified geology of the Mitchell and East Mitchell Zones, showing Stuhini Group undifferentiated sedimentary wall rock (SEDS), Hazelton Group volcano-sedimentary wall rock (VOLC), plagioclase-hornblende diorite porphyry (P2), the sheeted vein body with >50 vol.% quartz veinlets (P1; H1 hybrid zones hosted in wall rock), zones with 25-50 vol.% quartz veinlets (P1B; H1B where hosted in wall rock), syn-mineral plagioclase-hornblende-K-feldspar-phyric porphyry dykes (P3), late syn-mineral porphyry dykes with coarse K-feldspar (K3), monzonite to syenite porphyry (P8), intrusion breccias (P5), post-mineral dolerite dykes (K4), and overburden (OVBD). The traces of drill holes occurring within ± 50 m of the section are shown as black lines. b) Distribution of hydrothermal alteration assemblages within the Mitchell deposit, illustrating the extent of propylitic alteration (PRO), potassic alteration (POT1, POT2), hornfels (HF), hornfels with biotite (HFB), hornfels with tourmaline (HFT), sericitic alteration (QSP), advanced argillic alteration (ADA), as well as the occurrence of hydrothermal anhydrite (ANHY). c) Contours of logged volume percent of total quartz veins within Mitchell drill core. d) Block model Au grades, showing clear concentric zoning; e) block model Cu grades, showing a concentric zoning pattern like that of Au; and f) block model Mo grades.

 

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7.2.4East Mitchell Zone

 

As discussed in Section 7.2.3, the East Mitchell Zone, found in the hanging wall of the MTF, is the displaced upper portion of the Mitchell Zone. Based on lithological, alteration, and metal zonation, the displacement on the MTF is estimated at approximately 1,600 meters to the south-southeast. The East Mitchell Zone is exposed on a bluff to the east-southeast of the Mitchell deposit. The dimensions of the East Mitchell Zone resource area are approximately 1,400 m long, from the NNW to the SSE, and approximately 1,000 m wide, from the WSW to the ENE. The deposit has been recently deglaciated and forms a distinctive gossanous zone at surface, due to active oxidation of sulphides,

 

The East Mitchell Zone features the vertical continuation of lithological units, hydrothermal alteration assemblages, and mineralization found in the Mitchell Zone. Therefore, mineralization at East Mitchell is also genetically and spatially associated with the same Early Jurassic Sulphurets (Texas Creek) suite intermediate composition porphyry intrusions found in the Mitchell Zone (see Section 7.2.3). These intrusions are hosted in Upper Stuhini Group sedimentary strata, including sandstones, siltstones, mudstones, and conglomerates, as well as Lower Jurassic Hazelton Group andesitic volcaniclastic units belonging to the Jack Formation. The unconformity between the Stuhini and Hazelton Groups dips towards the north in the region surrounding East Mitchell.

 

The East Mitchell Zone is centred on diorite intrusions with high volumes of quartz veinlets that represent the upward continuation of syn-mineral diorite intrusions from the Mitchell Zone (see Section 7.2.3). The zone of intense quartz and sulphide veining (“P1” or “Sheeted Quartz Vein Zone” with over 50 volume percent quartz veins) observed in the Mitchell Zone continues and widens in the East Mitchell Zone. Gold and copper mineralization within the intrusions are associated with sulphides, notably chalcopyrite, and controlled by quartz stockwork or sheeted quartz veinlets. Whereas the mineralization in the Mitchell Zone is found almost exclusively within hydrothermally altered intrusive rocks, an important part of the mineralization at East Mitchell is contained within country rocks peripheral to syn-mineral intrusions. In particular, the volcaniclastic Lower Hazelton Group wall rock immediately adjacent to the central East Mitchell intrusions feature significant gold, copper, and molybdenum mineralization. In general, metal grades within the country rocks gradually diminish at increasing lateral distances from the central, syn-mineral intrusions. However, at the southern end of the East Mitchell Zone, an area historically referred to as the “Gold Zone” is hosted within Lower Hazelton andesitic volcanic breccias and lapilli tuffs and features the highest gold grades of the deposit. The “Gold Zone” contains little quartz veining or copper sulphides but is correlated with anomalous zinc mineralization (e.g., >500 ppm Zn). The style of gold occurrence within the “Gold Zone” may represent the eroded remnants of a distinct hydrothermal event.

 

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Hydrothermal alteration assemblages observed at the East Mitchell Zone are characteristic of shallow to mid-level calc-alkaline porphyry Au-Cu-Mo systems – consistent with East Mitchell representing the displaced upper portion of the Mitchell Zone. Primary potassic alteration, characterized by an assemblage of chloritized biotite ± magnetite, is widespread in mineralized zones. Wall rocks typically retain the texture of hornfels. Extensive secondary quartz-sericite-pyrite (“QSP”) phyllic alteration overprints much of the East Mitchell Zone, apart from the deep, southern parts of the deposit. Locally in the northern and central parts of East Mitchell, QSP alteration grades into advanced argillic (“ADA”) alteration, characterized by the presence of the clay mineral pyrophyllite. The transition from white mica (muscovite)-bearing QSP alteration assemblage to a pyrophyllite-bearing assemblage is difficult to determine with the naked eye but can be detected using instrumentation such as a handheld shortwave infrared spectrometer. Discrete zones containing small volumes of anhydrite veinlets are found near the base as well as on the eastern flank of East Mitchell. Whereas tourmaline is found only in small, discrete areas in the shallow, eastern part of the East Mitchell Zone, a tourmaline-bearing zone, approximately 200 m wide, is found along the eastern side of East Mitchell. This tourmaline-bearing zone is associated with locally elevated copper grades.

 

As is the case with the Mitchell Zone, East Mitchell underwent significant ductile deformation during the mid-Cretaceous development of the Skeena fold and thrust belt. Penetrative cleavage is observed in most areas with QSP or advanced argillic alteration. East Mitchell is bounded at depth by the MTF, and along its eastern periphery by the northerly-trending Brucejack fault.

 

The East Mitchell Zone has been impacted by the Snowfield Landslide, which is an actively displacing area of bedrock slope instability, a response to the retreat of the Mitchell Glacier buttressing effects and accelerated weathering due to oxidation of pyritic, fractured, phyllic altered rocks. It affects much of the East Mitchell Zone and covers an area of approximately 900 by 1400m to a depth of up to 170m, slumping towards the Mitchell Valley. Displacement is at most a few tens of meters, and abundant oxidized fractures and clay seams or slip surfaces are observed within, and largely restricted to, this landslide zone.

 

A geological map of the East Mitchell Zone is shown in Figure 7.8. A representative cross section and 950 masl level plan of the Mitchell and East Mitchell Zones, showing lithological units, alteration assemblages, quartz vein abundance contours, gold grade, copper grade, and molybdenum grade, are shown in Figure 7.9 and Figure 7.10, respectively. An additional cross section of the East Mitchell Zone, intersecting the “Gold Zone,” is found in Figure 7.11.

 

 

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Figure 7.11Vertical section through the East Mitchell Zone (Seabridge, 2022)

 

 

Notes:Location is shown as B-B’ in Figure 7.10a. The traces of the Mitchell thrust fault (MTF) and Snowfield landslide (LAN) surface are shown on each panel. a) Simplified geology of the East Mitchell Zone, showing Stuhini Group undifferentiated sedimentary wall rock (SEDS), Hazelton Group volcano-sedimentary wall rock (VOLC), plagioclase-hornblende diorite porphyry (P2), the sheeted vein body with >50 vol.% quartz veinlets (P1; H1 hybrid zones hosted in wall rock), zones with 25-50 vol.% quartz veinlets (P1B; H1B where hosted in wall rock, post-mineral dolerite dykes (K4), and overburden (OVBD). The traces of drill holes within ± 50 m of the section are shown as black lines. b) Distribution of hydrothermal alteration assemblages within the Mitchell deposit, illustrating the extent of propylitic alteration (PRO), potassic alteration (POT1), hornfels (HF), hornfels with tourmaline (HFT), sericitic alteration (QSP), advanced argillic alteration (ADA), as well as the occurrence of hydrothermal anhydrite (ANHY). c) Contours of logged volume percent of total quartz veins within Mitchell drill core. d) Block model Au grades, with the elevated gold content of the “Gold Zone” visible in the upper right part of the panel; e) block model Cu grades; and f) block model Mo grades.

 

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7.2.5Iron Cap Zone

 

The Iron Cap deposit is the northernmost porphyry gold-copper-molybdenum deposit in the KSM district, and structurally above the Mitchell deposit, in the panel of rocks between the MTF and STF. It is now known to be hosted by an Early Jurassic intrusive complex that is roughly contemporaneous with those in the Mitchell and Kerr zones. The Iron Cap mineralized zone forms a tabular body striking roughly north-south and dipping ~60° to the west, with approximate true dimensions of 1,500 m in strike, 1,500 m down dip, and up to 800 m in thickness. Mineralization remains open down dip.

 

Only the southern and eastern margins of the Iron Cap deposit outcrop at surface, where weak to moderate gold + copper ± molybdenum mineralization is primarily hosted in sedimentary strata of the Hazelton Group. It is now known that Iron Cap glacier covers the central and northern portions of the deposit, and that the western side of the deposit is concealed beneath the largely barren hanging wall of the STF. Drilling campaigns during the 2016, 2017, and 2018 field seasons, which included drill holes that collared on the glacier and on the ridge to the west of Iron Cap, significantly expanded the known resource at Iron Cap.

 

The Early Jurassic Iron Cap intrusive complex is composed of multiple intrusion and breccia phases. One of the earliest phases is a pre-mineral, medium-grained plagioclase-hornblende-phyric diorite with porphyritic texture found on the southeast side of the complex (“P2 East”). A second plagioclase-hornblende-phyric diorite phase, finer-grained than the pre-mineral diorite, is northwest of the complex (“P2 West”). The P2 West phase is thought to be syn-mineral, as it hosts significant volumes of early quartz-sulphide veins (A-type and B-type veins, based on the vein classification scheme of Gustafson and Hunt, 1975), and because this diorite phase is spatially associated with some of the highest gold and copper grades observed at Iron Cap. Syn-mineral, plagioclase-K-feldspar-hornblende(-quartz)-phyric monzonite intrusions (“P3 East”) are in the central to eastern parts of the Iron Cap intrusive complex. The P3 East monzonites are typically medium-grained, with seriate texture, and include several thin dykes as well as a ~100 m thick tabular intrusion striking roughly north through the centre of the Iron Cap deposit, and dipping ~60° west. As with the syn-mineral diorite, portions of the P3 East monzonite host significant volumes of early quartz-sulphide A-type and B-type veins and strong gold-copper mineralization. Finally, a suite of weakly-mineralized monzonite dykes and intrusions, coarser-grained than the central syn-mineral monzonite, occurs on the western side of the intrusive complex (“P3 West”). A radiometric U-Pb zircon age of 195.4 ± 2.1 Ma was obtained for the P3 East monzonite, which is in line with the other radiometric ages obtained for porphyry emplacement and mineralization within the Sulphurets district (190.3 ± 0.8 Ma to 197 ± 3 Ma; Bridge, 1993; Margolis, 1993; Febbo et al., 2019a; Febbo et al., 2019b). Unfortunately, U-Pb zircon age dating is not possible for the diorite intrusive phases at Iron Cap, as they do not contain zircon as an accessory mineral. The relative ages of the P2 West, P3 East, and P3 West intrusions are also unclear, due to the lack of cross-cutting relationships, but they are assumed to be roughly contemporaneous.

 

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Mineralized hydrothermal breccias are significantly more abundant and voluminous at Iron Cap than at the Mitchell, Sulphurets, or Kerr deposits. Tabular, steeply dipping, and northerly striking mineralized hydrothermal breccia bodies, ranging from ~10 m to ~250 m wide, occur within central Iron Cap. The breccias are matrix-supported and contain poorly-sorted subangular clasts. The clasts within the breccias are commonly too altered to permit protolith identification but clasts of P3 monzonite are sometimes observed. Jigsaw hydrothermal breccias are also common along the margins of the P2 East intrusion, with subangular to angular monomict clasts of the P2 East intrusion in a hydrothermally-altered rock flour matrix.

 

Post-mineral intrusions in the Iron Cap intrusive complex are limited to a small number of volumetrically insignificant dykes. K-feldspar-phyric monzodiorite dykes have been identified in the northern half of Iron Cap. These plagioclase-K-feldspar-hornblende-phyric dykes contain plagioclase and hornblende phenocrysts up to 8 mm and K-feldspar phenocrysts up to 2 cm, and closely resemble the post-mineral K-feldspar megacrystic dykes observed at the Kerr deposit. In the Sulphurets district, these dykes are thought to represent the last pulse of the Early Jurassic porphyry magmatic system. Rare aphanitic, mafic to ultramafic, dark green to blackish post-mineral dykes with meter-scale thicknesses also occur. These dykes are common throughout the rest of the Sulphurets district.

 

The Iron Cap deposit features hydrothermal alteration assemblages typical of porphyry Cu-Au deposits, including deep, central potassic alteration with an assemblage of K-feldspar – magnetite ± biotite (now chloritized), weak peripheral propylitic alteration with an assemblage of chlorite – epidote – carbonate ± hematite ± magnetite, and sericitic alteration. Potassic alteration is most commonly observed in the P2 West, P3 East and P3 West intrusions, as well as the deep hydrothermal breccias. Sericitic alteration zones, featuring an assemblage of fine-grained muscovite or illite – quartz – pyrite, are concentrated in the upper half of the deposit as well as within the hydrothermal breccias. The intensity of the sericitic alteration varies from the replacement of felsic minerals, to the partial replacement of mafic minerals, to the complete and texturally destructive replacement of all primary felsic and mafic phenocrysts. Strong sericitic alteration obliterates most original textures, making protolith identification difficult in places. Though present in minor quantities, the greatest abundances of galena, sphalerite, and silver or arsenic sulphosalts are found in the upper half of the Iron Cap Zone, and may be a consequence of telescoping or downward migration of the hydrothermal system. Hydrothermal anhydrite, which occurs abundantly in certain parts of the Kerr and Mitchell deposits, is conspicuously absent at Iron Cap.

 

Total quartz vein abundances at Iron Cap typically range from <1 vol.% to ~5 vol.%. However, a handful of discrete, narrow zones at Iron Cap feature notably high (>10 vol.%) abundances of sheeted A-veins, which correspond to zones of particularly elevated gold and copper grades. Such zones are notably observed within the P3 East and P2 West intrusive phase. The vein types present at Iron Cap include: early quartz – chalcopyrite ± magnetite “A-veins”, magnetite ± chalcopyrite veins with selvedges of pink hydrothermal K-feldspar ± magnetite; quartz – chalcopyrite ± molybdenite ± pyrite “B-veins,” with a distinctive sulphide-bearing central suture surrounded by quartz; gray quartz veinlets with a variable sulphide-sulphosalt assemblage of pyrite ± tennantite ± tetrahedrite ± chalcopyrite ± sphalerite, which crosscut the earlier A-veins and B-veins; pyritic “D-veins” with sericite-pyrite selvedges; and post-mineral quartz – carbonate ± chlorite ± sphalerite ± galena ± chalcopyrite ± tennantite ± tetrahedrite veins, which are thought to be associated with mid-Cretaceous deformation. High silver values are generally associated with presence of galena and sphalerite.

 

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Microscopic examinations of polished thin sections confirm that Iron Cap was also subjected to post-mineral deformation, as evidenced by widespread mylonitic textures.

 

The Iron Cap Zone terminates at the south along the north-dipping Iron Cap Fault (ICF). South of the fault, hornfelsed sedimentary rocks are mineralized with marginal gold and copper grades similar to intervals above the MTF at Mitchell. A few holes through this area contain higher than average molybdenum grades, including in interval of 133 m with 0.10% molybdenum.

 

A geological map of the deposit is presented in Figure 7.12. A representative cross section and 1200 masl level plan of the Iron Cap deposit, showing lithological units, alteration assemblages, quartz vein abundance contours, gold grade, copper grade, and molybdenum grade, are shown in Figure 7.13 and Figure 7.14, respectively.

 

Figure 7.12Iron Cap geology map. The traces of the Sulphurets Thrust Fault (STF), Johnstone Fault (JF), and Iron Cap Fault (ICF) are shown. (Seabridge, 2019)

 

 

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Figure 7. 13Vertical cross-section through the Iron Cap deposit, looking NNE (Seabridge, 2019)

 

 

Notes:Location of section is shown on Figure 7.12. The traces of the Sulphurets Thrust Fault (STF) and the Iron Cap Fault (ICF) are shown. a) Simplified geology of the deposit, showing undifferentiated wall rock (WR) and the principal intrusions and breccias of the Iron Cap intrusive complex: P2 East (P2E), P3 West (P3W), P3 East (P3E), and hydrothermal breccias (BX). See text for unit descriptions. The traces of drill holes within ±100 m of the section are shown as black lines. b) Hydrothermal alteration zoning illustrating the extent of potassic (POT.), moderate sericitic (MOD. SER.) and strong sericitic (STR. SER.) alteration assemblages (see text), as well as the extent of arsenic concentrations >50 ppm; c) contours of logged volume percent of total quartz veins within drill core; d) Au grades from the Seabridge Gold EOY 2018 block model; e) Cu grades from the Seabridge Gold EOY 2018 block model; and f) Mo grades from the Seabridge Gold EOY 2018 block model.

 

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Figure 7. 14Plan view through the Iron Cap deposit at 1200 m elevation (Seabridge, 2019)

 

 

Notes:The traces of the Sulphurets Thrust Fault (STF), Johnstone Fault (JF), and Iron Cap Fault (ICF) are shown. a) Simplified geology of the deposit, showing undifferentiated wall rock (WR) and the principal intrusions and breccias of the Iron Cap intrusive complex: P2 East (P2E), P2 West (P2W), P3 West (P3W), P3 East (P3E), hydrothermal breccias (BX) and post-mineral K-feldspar megacrystic dykes (P4). See text for unit descriptions. The traces of drill holes within ±100 m of the section are shown as black lines. b) Hydrothermal alteration zoning illustrating the extent of potassic (POT.), moderate sericitic (MOD. SER.) and strong sericitic (STR. SER.) alteration assemblages (see text), as well as the extent of arsenic concentrations >50 ppm; c) contours of logged volume percent of total quartz veins within drill core; d) Au grades from the Seabridge Gold EOY 2018 block model; e) Cu grades from the Seabridge Gold EOY 2018 block model; and f) Mo grades from the Seabridge Gold EOY 2018 block model.

 

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8.0Deposit Types

 

Porphyry copper ± gold ± molybdenum systems are defined as large volumes (~10 km3 – 100 km3) of mineralized and hydrothermally altered rocks centered on porphyry intrusions (Sillitoe, 2010). Porphyry systems are generated at convergent plate margins and along magmatic arcs above subduction zones. These systems typically span the uppermost ~4 km of the crust (Singer et al., 2008), with their stocks and dykes connecting downwards to deeper parental magma chambers that double as the source of high-temperature, high-pressure metalliferous fluids for the system (e.g., Cloos, 2001). The parental magmas are water-rich, highly oxidized, and exceptionally sulphur-enriched, and range in composition from granite to diorite to monzonite (e.g., Sillitoe, 2010). A metal-rich aqueous phase is released from cupolas atop the parental chambers during cooling and fractionation, and is episodically transported upward along with porphyry dykes. As they rise, the magmas and fluids undergo drastic temperature, pressure, and chemical changes that result in the precipitation of metals and the formation of distinctive hydrothermal mineral assemblages in an upward and outwardly zoned pattern.

 

The KSM deposits display many diagnostic features of porphyry Cu ± Au ± Mo systems. The deposits are centered on intrusive complexes composed of Early Jurassic, Texas Creek suite porphyry stocks and dykes. The Kerr, Mitchell, East Mitchell, and Iron Cap zones display the typical lateral and vertical zoning sequence of alteration assemblages observed in many porphyry systems: deep central potassic alteration, peripheral propylitic alteration, and shallow sericitic alteration. Mineralization is associated with stockwork quartz veinlets and arrays of sheeted quartz veinlets, with vein density decreasing in later intrusive phases. Host rocks may be mineralized for up to several hundred meters from the intrusions. The Kerr, Mitchell, and East Mitchell zones also feature small remnants of advanced argillic alteration. The structurally complex Sulphurets deposit does not feature the same clear alteration zoning patterns observed at the other three deposits, due to its dismembered and fragmental nature. However, the Lower Panel fault block at Sulphurets, which hosts the bulk of the mineralization, features potassic alteration and mineralization typical of porphyry Cu ± Au ± Mo systems.

 

Other large gold-enriched porphyry copper deposits, comparable to those in the KSM district, include Pebble (Alaska), Bingham (Utah), Grasberg (Indonesia), and Galore Creek (British Columbia). The KSM district porphyry Au-Cu-Mo deposits feature large volumes of plagioclase-hornblende-phyric diorite that is intimately associated with mineralization and smaller volumes of more evolved magmas, including granodiorite and monzodiorite dykes. Diorite is at the mafic end of the compositional spectrum of magmas commonly associated with porphyry Cu ± Au ± Mo formation. Another example of a gold-enriched porphyry copper deposit with syn-mineral diorite intrusions is Cerro Casale, Chile (e.g., Palacios et al., 2001).

 

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Porphyry Cu ± Au ± Mo systems are subdivided into two categories: calc-alkaline and alkaline Cu ± Au ± Mo deposits. The KSM district deposits mainly display features that are characteristic of calc-alkaline porphyry Cu ± Au ± Mo deposits: high quartz vein volumes, strong pyrite mineralization, areas of intense sericitic alteration, and an association with calc-alkaline intrusions. However, discrete zones within the district, restricted to fault panels that have been thrust over the main deposits, display characteristics consistent with alkalic style porphyry mineralization. These zones include the hanging wall of the Mitchell thrust fault above the main Mitchell deposit, as well as the Upper Panel fault block of the Sulphurets deposit in the hanging wall of the Sulphurets thrust fault. These alkalic-style zones contain largely barren porphyritic monzonite to syenite intrusions with potassic alteration and reddish hematite dusting, surrounded by volcano-sedimentary wall rock and hydrothermal breccias that contain weak to moderate copper-gold mineralization. The alkalic-style mineralized zones represent only a very minor portion of the total district resource.

 

The principal sulphides observed in the KSM deposits are pyrite and chalcopyrite, with minor molybdenite, and trace amounts of sphalerite, galena, tennantite-tetrahedrite, bornite, enargite, and arsenopyrite. Magnetite and hematitized magnetite are common, especially in deeper parts of the deposits, and hydrothermal anhydrite is observed in certain areas of the Kerr, Mitchell, and East Mitchell deposits. Native gold is rarely observed macroscopically, and is primarily detected as microscopic clusters at sulphide grain boundaries or as inclusions. All mineralization is hypogene, except for a small remnant of preserved supergene mineralization at the upper limits of the Kerr deposit where chalcocite coatings on pyrite and chalcopyrite have been observed, at the Main Copper (Sulphurets) occurrence where a remnant of leached capping and partial oxide mineralization is preserved at the highest elevations, and at East Mitchell within the Snowfield Landslide which has fracture controlled oxidation and local leaching in slumped intervals.

 

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9.0Exploration

 

This section summarizes Seabridge’s 2011 through 2021 non-drilling exploration programs at KSM. Since 2006, the majority of Seabridge’s exploration activities have been focused on various drilling programs that were designed to test recognized mineralized zones as well as several exploration targets. The following sections summarize by year various non-drilling exploration activities that Seabridge has completed.

 

9.12011 Geophysical Exploration Program

 

A Spartan MT (magneto telluric) survey was carried out by Quantec Geoscience Limited, over a period of 37 days from July 21, 2011 to August 26, 2011. A total of 175 MT Logger sites at an average spacing of 500 m were completed covering the KSM property from Kerr to Iron Cap zones and tunnel route north of the KSM grid. The data were collected over a frequency range of 320 Hz to 0.001 Hz with variable site spacing. The exploration objectives of the Spartan MT survey at the KSM Property were to detect mineralization and/or associated alteration zones, establish an understanding of the geological system and fluid pathways to great depth, and generate a 3D geophysical model by MT data to integrate with KSM geological model.

 

9.1.1Results of 2011 Geophysical Program

 

The 3D inversion results indicate that the subsurface resistivity, from the surface to a depth of approximately 2 km, varies over a range of 100 Ωm to 5000 Ωm. The southern half of the KSM property reveals a large conductive body with a thickness of more than 2 km, interpreted to be a large magmatic intrusion and favourable for porphyry system.

 

The central and northern parts of the KSM Property show relatively shallow conductive zones with thickness of approximately 500 m and are associated with alteration zones, which are the results of large scale thrust faulting in the area. The resistivity model also helped to map a number of sub-parallel faults and lineaments.

 

9.22013 Geophysical Exploration Program

 

SJ Geophysics Ltd. was contracted to acquire geophysical data in several boreholes at KSM for Seabridge. Borehole total magnetic field, induced polarization (IP), and time domain electromagnetic (TDEM) techniques utilizing SJ Geophysics’ borehole Volterra-EM system were conducted on five drill holes at the Kerr, McQuillan, and Iron Cap zones.

 

The objectives were to determine if the known deposit zones have an electromagnetic and/or magnetic signature and to identify if there is a relation between quantities of magnetite and gold-copper mineralization.

 

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9.2.1Results of 2013 Geophysical Program

 

The EM data did not identify any strong conductors in the holes surveyed. Based on the current holes surveyed, there does not appear to be an EM response to the deposit. In borehole K-13-24, a weak off-hole EM anomaly was identified that is believed to correlate with a low resistivity zone. Unfortunately, IP data were not collected in this hole, so this cannot be verified. In drill hole K-13-25, chargeability was found to increase with depth, and more chargeable rocks are located towards the east of the drill hole.

 

In borehole MQ-13-06, the total field magnetic data show three major magnetic units are present. Two off-hole EM anomalies were identified, with the larger one being located at approximately 850 m and to the southeast of the drill hole, mostly likely within 50 m to 100 m. Again, this conductor was poorly coupled. It may be better to use smaller, more focused loops in the future.

 

In the Iron Cap zone, boreholes IC-13-48 and IC-13-49 showed very good agreement between each other. From the magnetic data, the boundaries of a relatively constant magnetic unit were identified. This unit correlates with a low resistivity and high chargeability unit based on the IP data. It is believed that this may be a mineralized zone. The IP data from IC-13-48 indicates that more chargeable rocks are located towards the east and south of the drill hole.

 

S.J.V. Consultants Ltd. (SJV) was retained to review and interpret all historical geophysical data with the intention of identifying characteristic signatures that might be related to observed mineralization and geology. The IP data were reviewed and input to the UBC 2D resistivity and IP inversion algorithms, with the aim of improving the available models by using more modern algorithms as well as a finer cell size. The output models show some significant differences from the previous results and also have significantly better resolution of features at depth, as a result of the finer cell size.

 

The Kerr deposit overlies a resistive body but is itself hosted within a moderately conductive zone, close to the background level. There are no available IP or magnetic data for this area. The Sulphurets deposit also sits on the margins of a resistive body. The more detailed IP resistivity identified a narrow, dipping conductive zone that correlates very well with the actual mineralization. The magnetic susceptibility in this area is moderate. The Mitchell deposit is located along a conductive feature that correlates well with Mitchell creek. This is interpreted to represent a fault zone that may have provided a pathway for mineralizing fluids and/or intrusions. Mitchell sits within a clear magnetic low, which is interpreted to represent magnetite-destructive alteration.

 

The Iron Cap deposit is located within an area of high chargeability, but was not well defined on either the MT or IP resistivity data. The mineralization sits on the boundary of a large magnetic high that has significant roots in the 3D inversion.

 

The most direct correlation between the geophysics and the mineralization occurred at the Mitchell deposit, which is hosted by a prominent magnetic low, interpreted to be a result of magnetite-destructive alteration. Other magnetic lows within the property are therefore considered high priority targets. The 2D IP provides better resolution of the resistivity at relatively shallow depths than the MT survey, which is focused on deep exploration. Interpretation of the IP was somewhat limited as the survey covers only the northern portions of the KSM area and the lines are spaced too far apart to allow good correlation of features between lines. The relationship of the mineralization to the IP resistivity and chargeability was variable and complicated. Low resistivity features within high chargeability zones could be correlated to the mineralization at the Sulphurets deposit and are therefore recommended as possible future targets.

 

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9.32014 Geophysical Exploration Program

 

Quantec Geoscience Ltd. was contracted to perform Spartan MT and AMT (audio magneto telluric), gravity, and IP surveys over the Property. The MT/AMT survey was designed to expand coverage of the 2011 survey and selectively infill areas at a higher resolution with readings at approximately 250 m spacings. Gravity readings were collected at the same survey sites. Two east-west lines of IP surveying were made, one in Sulphurets valley and one in Mitchell valley.

 

9.3.1Results of 2014 Geophysical Programs

 

The gravity survey utilized a L&R Model G743 Base station, and the Geosoft Reduction processing platform with density of 2.65 g/cc. Corrections were made for latitude, terrain (regional and local), free-air anomaly calculation, and bouguer anomaly calculation. There were a total of 154 readings unevenly distributed over an area of about 75 skm2. The strong correlation between high readings and high topography suggests the topography is too extreme and the model too coarse to resolve and interpret the data correctly.

 

The Spartan MT survey utilized the same equipment and parameters as the survey done in 2011. A total of 49 readings produced acceptable results, and 23 readings were rejected due to excessive “noise”. Attempts to obtain readings on permanent snow and ice were unsuccessful, so the area covered was not expanded significantly from 2011.

 

IP surveying was performed on east-west lines of about 5 km each in the bottom of Sulphurets and Mitchell valleys, to reach the minimum elevation penetration of the survey. In Sulphurets valley, a strong chargeability response occurs east of the trace of the Sulphurets thrust fault and reflects continuity of pyritic alteration between Kerr and Sulphurets deposit. Similarly, in Mitchell valley high chargeability responses occur east of the STF trace.

 

9.42015 Geophysical Exploration Program

 

Precision Geosurveys Inc. was contracted to run a helicopter supported magnetic and radiometric survey covering almost all of the KSM claims, to improve resolution of the subsurface geological model in undrilled areas. The magnetic sensors are flown in a non- magnetic and non-conductive nose stinger configuration with 3D compensation, which allows the survey to be safely flown at reduced terrain clearance to minimize noise, improve resolution, and reduce the need for complex corrections to the data. The survey covered an area of about 1,600 km2. at a flight line spacing of 100 m. KSM mineralized areas are reflected by low susceptibilities where overprinting phyllic alteration has converted magnetite to pyrite, and high susceptibilities where potassic alteration with magnetite is still preserved.

 

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9.52019 Geophysical Program

 

During 2019, Geotech Ltd. carried out a helicopter-borne geophysical survey which included a caesium Z-axis tipper magnetometer, which is known as a Z-Axis Tipper Electromagnetic (ZTEM) system, a global positioning system (GPS) navigation system and a radar altimeter. In a ZTEM survey, a single vertical-dipole air-core receiver coil is flown over the survey area in a grid pattern, similar to regional airborne EM surveys. Two orthogonal, ferrite-core horizontal sensors are placed close to the survey site to measure the horizontal EM reference fields. Data from the three sensors are used to obtain the Tzx and Tzy tipper components at six frequencies in the 30 Hz to 720 Hz band. The ZTEM is useful in mapping geology using resistivity contrasts and magnetometer data provides additional information on geology using magnetic susceptibility contrasts. The survey covered an area of 190 km2. with a total of 1,000 line km flown at a 200 m spacing in an east-west direction. The survey was carried out between July 25 to September 8, with 23 days lost to weather, 10 due to equipment malfunctions, and 9 survey days flown.

 

In July and August, Dias Geophysical Ltd. was contracted to perform an IP survey over part of the Sulphurets zone area extending westward, to assist in mapping sulphide distributions and orientations in the area of drill testing, covering 3 km2. The survey utilized a continuously rolling distributed array pole-dipole using the common voltage reference (CVR) method, using a 200 m line spacing and a receiver spacing of 100 m. This approach has inherent flexibility in calculating multi-scale dipoles from the acquired data. The survey data can be processed in the conventional fashion, in which 50 m dipoles are calculated for the entire survey. If the signal to noise ratio is found to be too low the data can be reprocessed to increase the dipole separation to 200 m or 400 m to increase the ratio. The southeast corner of the survey covered a portion of the Sulphurets zone beneath the STF, and high chargeability features correlate very well with the modelled sulphide distribution previously determined from drilling. The exploration area targeted in 2019 above the STF displayed more limited and discontinuous high chargeability areas, which drilling confirmed were due to high concentrations of pyrite with minor chalcopyrite in veins, disseminations and patchy replacements associated with skarn alteration near intrusive/wall rock contact zones.

 

Mira Geoscience was contracted to undertake a regional district scale integrated model and targeting exercise for KSM, including the compilation of all digital exploration data and construction of an integrated 3D geological and structural model, as a basis for constrained inversion of magnetics, MT and IP surveys data. Modelling is based on available historical data, literature, geology, structures, geochemistry and geophysics datasets provided by Seabridge. The integrated 3D model was used to produce a set of targets based on a porphyry deposit concept involving both knowledge driven, and data driven (machine learning) analysis of the data and model.

 

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Deep penetrating geophysical techniques were employed to improve resolution on targets and generate discrete zones for testing in the future. New ZTEM surveys and 3D IP surveys were completed. Data was integrated into a digital 3D earth model by Mira Geoscience. These results are now being integrated with historical MT surveys, airborne high-resolution magnetic survey, bore hole geophysical surveys and geological mapping. Geophysical profiles indicate that these targets can be tested from surface but would likely be evaluated as bulk underground opportunities.

 

9.6East Mitchell Exploration

 

Between 1960 and 1983 Granduc Mines and Esso Minerals, the first modern mineral explorers in the area, carried out regional scale reconnaissance prospecting. This work led to geological mapping, rock sampling focused on the East Mitchell zone. In 1983 Esso Minerals mucked-out, surveyed and sampled totaling 126 meters in fifteen trenches yielding 64 chip samples. Other than a low resolution, regional airborne magnetic survey conducted by Newmont in the early 1960’s, no geophysical surveys are documented to have been undertaken. This is probably due to the extensive exposure over much of the area allowing rock sampling and mapping to adequately define drill hole targets. A total of 4 assessment reports of exploration work were filed between 1993 and 2008, mostly documenting diamond drilling.

 

The first drilling by Seabridge was done in 2021. Table 6.3 provides a summary of drilling campaigns and property ownership changes up to the present time.

 

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10.0Drilling

 

10.1Introduction

 

Drilling methods, procedures, extent of drilling, and relevant results for the Property have been described in previous NI 43-101 Technical Reports. Plan maps and representative drill hole cross sections through the mineral deposits are shown in Figures 7.2, 7.5, 7.8, 7.12, 14.1 to 14.4, 14.6 to 14.9, 14.11 to 14.14, and 14.20 to 14.23 to illustrate a summary and interpretation of the drilling results.

 

The majority of KSM drilling information that is stored in the end-of-year 2021 AcQuire database (which now includes the East Mitchell zone) was collected by Seabridge (67%). Seabridge has conducted annual drilling campaigns at KSM beginning in 2006. The remaining 33% of the drilling data were collected by Pretium and Silver Standard (24%), Placer (5%) and Falconbridge/Noranda (about 1%), with the remainder collected by five other companies (3%). The 2005 Falconbridge drill campaign was conducted while the Property was under option from Seabridge. The 2006 drill campaign was managed by Falconbridge under a joint venture (JV) with Seabridge. A summary of all KSM drill hole data organized by year is shown in Table 10.1. The majority of the 968 core holes shown in Table 10.1 were used to estimate Mineral Resources disclosed in this Report, but some of the data tested several non-resource targets in the KSM Property. Table 10.2 summarizes drilling data by company through 2021. The companies listed in Table 10.2 have been arranged in approximate chronological order starting with Esso Minerals in the 1960s. Minor core drilling was completed at KSM by several companies in the early 1980s, but ramped up significantly in the late 1980s and early 1990s by Placer Dome. Seabridge systematically added to the KSM drill hole data after their entry into the district in 2000, with annual drill campaigns beginning in 2006. The bulk of drilling at East Mitchell was done by Silver Standard Resources, predecessor of Pretium Gold, between 2006 and 2010.

 

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Table 10.1KSM Core Drilling Through 2021

 

 

Table 10.2Drilling by Company Through 2021

 

 

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Table 10.3 summarizes drilling at KSM by deposit and company through 2021. The data shown in Table 10.3 are the drill holes that were used to estimate Mineral Resources that are the subject of this Report.

 

Table 10.3KSM Drill Hole Summary by Area and Company Through 2021

 

 

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10.2Type and Extent of drilling

 

Seabridge has had long-standing relationships with local drilling and air support contractors that have allowed for a continually growing understanding about local drilling conditions. The drilling operations were conducted from the Sulphurets Creek camp which is located northwest of the Kerr deposit.

 

In 2019 Seabridge contracted with Driftwood Diamond Drilling out of Smithers, BC, to complete a 26-hole diamond core drilling program at the Sulphurets deposit.

 

No drilling was undertaken on the resource zones in 2020. In 2021, Hy-Tech Drilling from Smithers, BC, completed a program of infill drilling for metallurgical sampling, geotechnical testing, and resource confirmation at the Mitchell and East Mitchell zones. Helicopter support was provided by Summit Helicopters of Terrace, BC.

 

Figure 1.1 is a drill hole location map for the entire KSM district, showing all of the drilling data that were available to estimate Mineral Resources that are the subject of this Report (drilling through 2021). The drill holes are colour coded (blue represents non-Seabridge and red represents Seabridge drilling). Detailed drill hole location maps are presented in Figure 10.2 to Figure 10.5 for the Kerr, Sulphurets, Mitchell and East Mitchell, and Iron Cap deposits, respectively. Figure 10.2 to Figure 10.5 also show the outline of conceptual resource pits and resource block cave projected to surface, that constrain the Mineral Resources for the Property.

 

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Figure 10.1KSM Drill Hole Locations

 

 

Source: Seabridge 2022

 

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Figure 10.2Drill Hole Locations – Kerr Deposit

 

 

Source: Seabridge 2022

 

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Drilling at the Kerr deposit has identified a mineralized area measuring roughly 2,400 m north-south by 800 m east-west, and about 2,200 m vertically. The drill hole spacing in the upper open pit resource area is approximately 50 m to 75 m. Drill hole spacing through the block cave resource, which has been classified as nearly all Inferred material, ranges between 100 m to 200 m. Representative drill hole cross sections and level plans through the Kerr deposit are shown in Figures 14.1 to 14.4

 

Figure 10.3Drill Hole Locations – Sulphurets Deposit

 

 

Source: Seabridge 2022

 

Drilling at the Sulphurets deposit has identified a mineralized area measuring roughly 2,200 m northeast-southwest by 550 m northwest-southeast, and about 330 m vertically. The drill hole spacing in the open pit resource area ranges between 50 m to 75 m. Figures 14.6 to 14.9 show representative drill hole cross sections and level plans through the Sulphurets deposit.

 

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Figure 10.4Drill Hole Locations – Mitchell Deposit

 

 

Source: Seabridge 2022

 

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Drilling at the Mitchell deposit has identified a mineralized area measuring roughly 1,600 m east-west by 1,500 m down-dip, and 850 m thick. The drill hole spacing in the upper open pit resource area is approximately 75 m to 100 m. Drill hole spacing through the block cave resource, which has been classified predominantly as Inferred material, ranges between 100 m to 200 m. Representative drill hole cross sections and level plans are shown in Figures 14.11 to 14.14.

 

Figure 10.5Drill Hole Locations – Iron Cap Deposit

 

 

Source: Seabridge 2022

 

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Drilling at the Iron Cap deposit has identified a mineralized area measuring roughly 1,500 m northeast-southwest, by 1,500 m northwest-southeast, and about 850 m thick. The drill hole spacing in the upper block cave resource shapes ranges from 70 m to 75 m. Drill hole spacing through the lower block cave resource, which has been classified predominantly as Inferred material, ranges from 100 m to 200 m. Representative drill hole cross sections and level plans through the Iron Cap deposit are shown in Figures 14.22 to 14.25.

 

10.3Drilling Procedures

 

All of the drilling data collected in 2005 and between 2008 and 2018 was completed by one drill contractor, Hy-Tech Drilling Ltd., located in Smithers, BC. The drilling was completed using Hy-Tech’s Tech-5000 Fly Rigs utilizing HQ, NQ, BQ, and AQ rods. In 2006 drilling was completed by Boart Longyear, in 2007 by Blackhawk, Radius, and Driftwood drilling of Smithers, BC. Geotechnical drilling designated by hole number prefixes of KC-, MW-, and TS-, was completed by Cabo, Geotech, and Taltech Drilling, all based in BC. Nearly all of the helicopter support was provided by Lakelse Air Ltd. using Eurocopter A-Star machines. Other helicopter services were provided at various times by Mustang Helicopters and Summit Helicopters. Drilling operations were conducted from the Sulphurets Creek camp, which is located northwest of the Kerr deposit.

 

Seabridge used directional drilling and conventional wedging methods for a portion of some of their deeper drilling programs at Kerr, Iron Cap, and Mitchell. Tech Directional Services Inc. from Ontario, Canada, were contracted to provide directional drilling services using DeviDrill equipment. DeviDrill uses a steerable wireline core barrel that allows a “daughter” hole to be wedged off of a “mother” hole and vectored towards a target zone with reasonable accuracy. Small diameter core (AQ) was retrieved during the crucial turn away from the mother hole so minimal data was lost. Bearing and inclination data were collected using a miniature electronic single-shot survey tool (DeviTool PeeWee) that is designed to pass through the DeviDrill bit. Information regarding this drilling method can be found at http://www.techdirectional.com/. Drill holes with a letter designation after the hole number represent wedged drill holes that utilized the directional drilling method. A total of 38 daughter holes were wedged off mother holes totalling approximately 27,534 m.

 

Drill core was placed into wooden core boxes by the drill contractor at the rig and delivered twice daily by helicopter from the rigs to Seabridge’s Sulphurets Creek camp. An inventory of the core was completed by Seabridge geologists, which included a review of core condition, a check of run block depths, and generation of a quick down-hole lithological log.

 

The drill core was typically scanned for various base metal quantities using a Niton handheld x-ray fluorescence (XRF) analyzer prior to cleaning the core. Seabridge has determined that a factor of 2.0 to 2.2 times the Niton copper reading closely approximates the assayed copper content percentage. The Niton readings are primarily used to alert/train the logging geologist about approximate copper percentages in mineralized intersections. That data was written on the core with wax markers and are visible on core photos. Magnetic susceptibility was also recorded for each drill hole using a handheld device. The mag readings were exported from the device as .csv files.

 

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After cleaning, the core was logged for lithology, alteration, structure, and oxidation state onto paper logs by Seabridge geologists. That information was later entered into Microsoft Excel spreadsheets by each logger. Separate paper logs were used to capture geotechnical information like core recovery, rock quality designation (RQD), and fracture count. The geotechnical logs are based on data between core run blocks.

 

Assay samples were laid out by the logging geologist. Samples were primarily laid out in 2-m lengths, but were broken at distinct lithologic, alteration, or mineralization contacts. Likewise, samples were broken at core diameter changes. The sample data were hand recorded onto paper logs with hole name, from depth, to depth, and various sulphide mineral estimates.

 

Pieces of drill core (14 cm to 20 cm long) were marked for bulk density determination about every 100 m down-the-hole by the logging geologist by labelling the wooden core box with “SG”. Those small pieces were not cut for assay sample. Periodically, a contract employee weighed the core pieces in air and water so that a bulk density could be calculated.

 

Prior to sawing, the drill core was photographed using a digital camera. After all logging procedures were complete, the core boxes were moved to the core cutting facilities located adjacent to the core logging tents.

 

A summary of the interpretation of the Kerr drilling results are illustrated in Figures 7.2 to 7.4 and 14.1 to 14.4.

 

A summary of the interpretation of the Sulphurets drilling results are illustrated in Figures 7.5 to 7.7 and 14.6 to 14.9.

 

A summary of the interpretation of the Mitchell and East Mitchell drilling results are illustrated in Figures 7.8 to 7.11 and 14.11 to 14.14.

 

A summary of the interpretation of the Iron Cap drilling results are illustrated in Figures 7.12 to 7.14 and 14.20 to 14.23.

 

10.4Recent Drilling – Mitchell and East Mitchell

 

The drilling results reported in Table 10.4 represent 2,460 m of the 4,531 m program completed in 2021 to confirm model grades, obtain metallurgical sample material, and evaluate geotechnical characteristics for slope determination. Results from these holes were compared to what was predicted by the resource model and generally confirmatory (see Figure 14.19 and 14.20).

 

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Table 10.42021 Drill Hole Assay Results

 

Drill Hole ID Total Length From To Thickness Gold Grade Copper Grade Silver Grade
  (m) (m) (m) (m) (g/t) (%) (g/t)
Mitchell Deposit
M-21-154 319.6 2.7 319.6 316.9 0.66 0.19 2.9
including 2.7 78.1 75.4 0.89 0.22 3.3
East Mitchell Deposit
SF-21-08 362.1 4.3 362.1 357.8 1.06 0.04 1.4
including 4.3 186.5 182.2 1.70 0.04 1.8
including 34.0 80.0 46.0 2.55 0.05 2.0
SF-21-09 150.5 0.0 150.5 150.5 1.81 0.02 1.3
including 7.5 62.5 55.0 2.89 0.02 2.0
SF-21-10 90.4 1.5 90.4 88.9 2.80 0.02 1.8
SF-21-11 90.2 1.6 90.2 88.6 2.64 0.03 2.0
SF-21-12 400.5 0.0 400.5 400.5 0.97 0.20 2.3
including 164.5 219.9 55.4 1.37 0.24 2.4
East Mitchell Geotechnical Holes
SF-21-06 300.4 183.5 210.5 27.0 0.77 0.06 1.2
SF-21-07 595.8 No significant assays
SF-21-13 150.6 3.9 73.3 69.4 0.58 0.05 0.9
and 98.1 114.9 16.8 0.47 0.04 1.0

 

10.5QP Comments Regarding Drilling and Sampling Factors

 

In general, core recovery for the various KSM drilling campaigns was excellent, averaging approximately 97%.

 

No material drilling, sampling, or recovery issues were encountered within the mineralized portions of the other deposits within the KSM Property that were drill tested during the 2006 to 2021 campaigns.

 

In the opinion of the QP responsible for this section of this Report, there are no drilling or sampling factors that could materially impact the accuracy and reliability of the assay results associated with the 2006 to 2021 KSM drilling. Furthermore, the QP responsible for this section of this Report believes that the assays associated with the various KSM drilling campaigns are suitable to be used to estimate Mineral Resources.

 

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11.0Sample Preparation, Analysis and Security

 

11.1Introduction

 

This section summarizes sample preparation, analyses, and security of KSM drill samples through the 2021 drilling campaign. Prior Technical Reports provided sampling details and quality control results from various annual drilling programs. In this Technical Report, a summary of sampling and assaying over the past 12 years is provided. In general, Seabridge has employed relatively consistent sampling methods over the years with minor modifications over the past five years regarding some assaying protocols. Initially, Seabridge used Eco Tech as their primary assay laboratory from 2006 to 2011 when Eco Tech was bought out by ALS Chemex, who has acted as Seabridge’s primary assay laboratory since that time to the present. Over the years, Seabridge’s quality control protocols have included the submission of certified standard reference materials (SRMs or “standards”) blanks, and duplicate field samples. Typically, 5% to 10% of the assay pulps from the primary laboratory were submitted to a secondary accredited assay laboratory for check assay comparison purposes. The following sections will summarize key aspects of sample preparation, analyses, etc., covering two time periods, pre-2012 and post-2012, primarily to account for the change of primary laboratory in 2011.

 

11.2KSM Sample Preparation Methods and Procedures

 

11.2.1Statement on Sample Preparation Personnel

 

Labourers contracted from Tahltan Native Development Corporation conducted all initial sample preparation (sawing and bagging) and were trained by, and under the direct supervision of, geologists employed by Seabridge. Drill core and quality control samples were shipped to the primary assay laboratory’s preparation facility (either to Eco Tech’s preparation facility located in Stewart, BC, or ALS Canada’s preparation facility located in Terrace, BC, or other ALS facilities). The prepared samples were then shipped to the primary assay laboratory’s analytical facility (i.e., either Eco Tech’s laboratory located in Kamloops, BC, or ALS Canada’s laboratory located in North Vancouver, BC).

 

11.2.2Sample Preparation and Dispatch

 

Drill core, placed in wooden core boxes by the drilling contractor, was loaded into metal baskets and delivered twice daily by helicopter from each drill rig to Seabridge’s Sulphurets Creek camp facilities. After the core was delivered to the core logging tents, Seabridge geologists took an inventory of the core (a review of core condition, a check of run block depths, and a quick down-hole lithologic log was prepared). After the drill core had been completely logged, the sample starting/ending points were then laid out. A numbered sample tag was stapled to the wooden core box at the beginning of the sample run and the core was photographed. Core boxes were then moved to the core cutting facilities located adjacent to the core logging tents. Two tents containing four saws with 14-inch diamond impregnated blades designed for rock cutting were utilized for sawing the core longitudinally. The saws were mounted on secure wooden stands at waist height. The saw blades were cooled, cleaned, and lubricated with fresh, non-recirculated water during cutting. The saw operator placed boxes of uncut core on tables adjacent to the saws and cut each piece of core sequentially within each marked sample interval. The assay half of the sample was placed in a heavy-duty polythene bag along with the sample number tag, the remaining half was placed back in the core box. The outside of the polythene bags was marked with the sample number and then stapled shut.

 

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The polythene sample bags were inventoried and then approximately three or four polythene sample bags containing HQ diameter core (or five or six bags with NQ core) were placed into large polyweave (rice) shipping bags. Approximately 30 rice bags were placed into wooden crates (“totes”) and/or large poly bulk bag for shipment. If only a tote is used each rice bag receives a security tag, if a bulk bag is used a security tag is placed on the bulk bag. The totes were flown by helicopter to a landing pad near km 54 on the Eskay Creek Mine Road located behind a locked gate. The totes were then placed into a locked steel sea-going container that can hold approximately eight totes. Once or twice a week, Granmac Services from Stewart, BC, picked up the totes and delivered them to an ALS receiving facility located in either Stewart, BC (pre-2013), or Terrace, BC, or by Bandstra Transportation and delivered to ALS facilities in either Kamloops, BC, or Vancouver BC (post-2019), where the samples were logged into ALS’s system. From the ALS preparation laboratory, the samples were delivered to the ALS assay laboratory located in North Vancouver, BC, by either ALS personnel or other commercial carriers.

 

11.2.3Pre-2012 Analytical Procedures

 

At the Eco Tech facilities in Stewart, samples were sorted and dried (if necessary), crushed through a jaw crusher and cone or roll crusher to –10 mesh, then split through a Jones riffle until a –250 g sub-sample was achieved. The sub-sample was pulverized in a ring and puck pulverizer so that 95% of the material passed a –140 mesh screen, then rolled to homogenize. The resulting pulp sample was placed in a numbered paper envelope and securely packed in cardboard boxes. These boxes were shipped via Greyhound freight services to the Eco Tech facilities located in Kamloops, BC.

 

At the Eco Tech’s laboratory in Kamloops, a 30 g sample size was split out from the pulp envelope and then fire assayed using appropriate fluxes. The resultant doré bead was parted and then digested with aqua regia followed by an atomic absorption (AA) finish using a Perkin Elmer AA instrument. The lower limit of detection for gold is 0.03 g/t or 0.001 oz/t. For other metals, a multi-element inductively coupled plasma (ICP) analysis was completed. For this procedure, a 0.5 g sample was digested with 3 mL mixture of hydrogen chloride, nitric acid, and water at a ratio of 3:1:2 that contained beryllium, which acts as an internal standard, for 90 minutes in a water bath at 95°C. The sample was then diluted with 10 mL of water and analyzed on a Jarrell Ash ICP unit. Eco Tech’s ICP detection limits (lower and upper) are summarized in Table 11.1.

 

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Assay results were then collated by computer and were printed along with accompanying internal quality control data (repeats and standards). Results were printed on a laser printer and were faxed and/or mailed to appropriate Seabridge personnel. Appropriate standards and repeat samples were included on the data sheet.

 

Table 11.1ICP Detection Limits – Pre-2012 Data

 

Element Lower Upper   Element Lower Upper
Ag 0.2 ppm 100.0 ppm   Mo 1 ppm 10,000 ppm
Al 0.01% 10.00%   Na 0.01% 10.00%
As 5 ppm 10,000 ppm   Ni 1 ppm 10,000 ppm
Ba 5 ppm 10,000 ppm   P 10 ppm 10,000 ppm
Bi 5 ppm 10,000 ppm   Pb 2 ppm 10,000 ppm
Ca 0.01% 10.00%   Sb 5 ppm 10,000 ppm
Cd 1 ppm 10,000 ppm   Sn 20 ppm 10,000 ppm
Co 1 ppm 10,000 ppm   Sr 1 ppm 10,000 ppm
Cr 1 ppm 10,000 ppm   Ti 0.01% 10.00%
Cu 1 ppm 10,000 ppm   U 10 ppm 10,000 ppm
Fe 0.01% 10.00%   V 1 ppm 10,000 ppm
La 10 ppm 10,000 ppm   Y 1 ppm 10,000 ppm
Mg 0.01% 10.00%   Zn 1 ppm 10,000 ppm
Mn 1 ppm 10,000 ppm        

 

11.2.4Post-2012 Analytical Procedures

 

ALS served as Seabridge’s primary assay laboratory. ALS is a leading provider of assaying and analytical testing services for mining and mineral exploration companies and has no association or affiliation with Seabridge. All of ALS’s locations are International Organization for Standardization (ISO) 9001:2000 certified.

 

AcmeLabs served as a check assay laboratory for Seabridge. AcmeLabs is a leading geochemical and assaying facility and has no association or affiliation with Seabridge. In October 2011, AcmeLabs’ Vancouver, BC, facility received ISO/International Electrotechnical Commission (IEC) 17025:2005 accreditations from the Standards Council of Canada.

 

Actlabs served as a check assay laboratory for Seabridge in 2019 and 2021. Actlab is a leading geochemical and assaying facility and has no association or affiliation with Seabridge. In June 2014, Actlabs’ Kamloops, BC, facility received ISO/IEC 17025:2017 accreditations from the Standards Council of Canada.

 

At the ALS preparation facility located in Terrace (or Kamloops), BC, samples were sorted and dried (if necessary), and crushed through a jaw crusher and cone or roll crusher to 70% –2 mm using ALS protocol CRU-31. The crushed sample was then split using ALS protocol SPL-21 using a riffle splitter. A portion of the crushed sample was replaced into the polythene bag (coarse reject) and stored temporarily at the ALS facility. A portion of the crushed sample was then pulverized using ALS protocol PUL-31 using a ring and puck pulverizer, until approximately a 250 g sub-sample (pulp) was achieved with 85% passing 75 µm or better. The resulting pulp sample was placed in a numbered paper envelope and securely packed in cardboard boxes. These boxes were shipped by ALS to their assay facility located in North Vancouver, BC.

 

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At the ALS analytical laboratory located in North Vancouver, a 30 g sample was split out from the pulp envelope and then fire assayed using ALS protocol Au-AA23 using appropriate fluxes. The resultant doré bead was parted and then digested with aqua regia followed by an AA finish using an Agilent AA 240 Series instrument. The lower limit of detection for gold is 0.005 g/t.

 

For other metals, a multi-element ICP-atomic emission spectroscopy (AES) analysis was completed using ALS protocol ME-ICP41. For this procedure, an approximately 0.5 g sample was digested with aqua regia in a graphite heating block. After cooling, the resulting solution was diluted to 12.5 mL with de-ionized water, mixed, and analyzed by ICP-AES using an Agilent ICP 720/730-ES Series instrument. The analytical results were corrected for inter-element spectral interferences. ALS’s ME-ICP41 lower and upper detection limits are summarized in Table 11.2 for the elements that Seabridge requested.

 

In 2019, a multi-element ICP-AES analysis was conducted using ASL protocol ME-MS41 (ultra-low detection limit mass spectroscopy). A prepared sample (0.50 g) is digested with aqua regia in a graphite heating block. After cooling, the resulting solution is diluted to with deionized water, mixed, and analyzed by inductively coupled plasma-atomic emission spectrometry. Following this analysis, the results are reviewed for high concentrations of bismuth, mercury, molybdenum, silver, and tungsten and diluted accordingly. Samples are then analyzed by ICP-MS for the remaining suite of elements. The analytical results are corrected for inter element spectral interferences. ME-MS21 lower and upper detection limits are summarized in Table 11.3.

 

In 2021, in addition to method ME-ICP41 (detailed above), a multi-element ICP- AES analysis was conducted using ASL protocol ME-ICP61. A prepared sample (0.25 g) is digested with perchloric, nitric, hydrofluoric, and hydrochloric acids. The residue is topped up with dilute hydrochloric acid and the resulting solution is analyzed by inductively coupled plasma-atomic emission spectrometry. Results are corrected for spectral interelement interferences. ME-ICP61 lower and upper detection limits are summarized in Table 11.4.

 

Table 11.2ICP Detection Limits – Post 2012

 

Element Units Lower Upper   Element Units Lower Upper
Ag ppm 0.2 100   Mo ppm 1 10,000
Al % 0.01 25   Na % 0.01 10
As ppm 2 10,000   Ni ppm 1 1,000
B ppm 10 10,000   P ppm 10 1,000
Ba ppm 10 10,000   Pb ppm 2 1,000
Be ppm 0.5 1,000   S % 0.01 10
Bi ppm 2 10,000   Sb ppm 2 1,000

 

table continues…

 

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Element Units Lower Upper   Element Units Lower Upper
Ca % 0.01 25   Sc ppm 1 1,000
Cd ppm 0.5 1,000   Se ppm 10 10,000
Co ppm 1 10,000   Sn ppm 10 10,000
Cr ppm 1 10,000   Sr ppm 1 1,000
Cu ppm 1 10,000   Th ppm 20 1,000
Fe % 0.01 50   Ti % 0.01 10
Ga ppm 10 10,000   U ppm 10 1,000
K % 0.01 10   V ppm 1 1,000
La ppm 10 10,000   W ppm 10 1,000
Mg % 0.01 25   Zn ppm 2 1,000
Mn ppm 5 50,000          

 

Table 11.3Aqua regia digestion, ICP-MS and ICP-AES analysis – 2019

 

Element Unit LDL UDL   Element Unit LDL UDL
Ag ppm 0.01 100   Mo ppm 0.05 10000
Al % 0.01 25   Na % 0.01 10
As ppm 0.1 10000   Nb ppm 0.05 500
Au ppm 0.02 25   Ni ppm 0.2 10000
B ppm 10 10000   P ppm 10 10000
Ba ppm 10 10000   Pb ppm 0.2 10000
Be ppm 0.05 1000   Rb ppm 0.1 10000
Bi ppm 0.01 10000   Re ppm 0.001 50
Ca % 0.01 25   S % 0.01 10
Cd ppm 0.01 1000   Sb ppm 0.05 10000
Ce ppm 0.02 500   Sc ppm 0.1 10000
Co ppm 0.1 10000   Se ppm 0.2 1000
Cr ppm 1 10000   Sn ppm 0.2 500
Cs ppm 0.05 500   Sr ppm 0.2 10000
Cu ppm 0.2 10000   Ta ppm 0.01 500
Fe % 0.01 50   Te ppm 0.01 500
Ga ppm 0.05 10000   Th ppm 0.2 10000
Ge ppm 0.05 500   Ti % 0.005 10
Hf ppm 0.02 500   Tl ppm 0.02 10000
Hg ppm 0.01 10000   U ppm 0.05 10000
In ppm 0.005 500   V ppm 1 10000
K % 0.01 10   W ppm 0.05 10000
La ppm 0.2 10000   Y ppm 0.05 500
Li ppm 0.1 10000   Zn ppm 2 10000
Mg % 0.01 25   Zr ppm 0.5 500
Mn ppm 5 50000          

 

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Table 11.44-acid detection limits by ICP-AES – 2021

 

1 Unit LDL UDL   Element Unit LDL UDL
Ag ppm 0.2 100   Mg % 0.01 25
Al % 0.01 25   Mn ppm 5 50000
As ppm 2 10000   Mo ppm 1 10000
B ppm 10 10000   Na % 0.01 10
Ba ppm 10 10000   Ni ppm 1 10000
Be ppm 0.5 1000   P ppm 10 10000
Bi ppm 2 10000   Pb ppm 2 10000
Ca % 0.01 25   S % 0.01 10
Cd ppm 0.5 1000   Sb ppm 2 10000
Co ppm 1 10000   Sc ppm 1 10000
Cr ppm 1 10000   Sr ppm 1 10000
Cu ppm 1 10000   Th ppm 20 10000
Fe % 0.01 50   Ti % 0.01 10
Ga ppm 10 10000   Tl ppm 10 10000
Hg ppm 1 10000   U ppm 10 10000
K % 0.01 10   V ppm 1 10000
La ppm 10 10000   W ppm 10 10000
Li ppm 10 10000   Zn ppm 2 10000
Mg % 0.01 25          

 

11.3Summary of the Nature, Extent, and Results of Quality Control Procedures

 

Since 2006, Seabridge has routinely submitted certified standard reference materials (“standards”) and barren (“blanks”) with their sawn drill core samples at a frequency rate of approximately 1 standard and 1 blank for every 32 to 33 regular samples. At the conclusion of each drilling campaign, about 5% to 10% of the pulps from the primary laboratory were sent to a secondary accredited assay facility providing an independent check of the original pulps. Starting in 2007, field duplicate drill core (sawn ¼ core) samples were collected and sent to the primary assay laboratory at a frequency of about one duplicate sample for every 54 regular samples. Table 11.5 summarizes the number of control samples that were submitted by Seabridge on an annual basis.

 

Table 11.5Summary of KSM Control Samples Submitted Thru Time

 

Year Regular
Samples
Submitted
Number of QA/QC Samples Submitted
Standards (S) Blanks
(B)
Duplicates
(D)
Pulp Checks 1 SBD Checks 2
2006 4,480 148 138 0 316 0
2007 7,685 243 236 125 760 53
2008 8,616 265 262 135 879 63
2009 6,338 195 192 96 580 30
2010 13,556 386 380 182 1,323 79
        table continues…

 

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Year Regular
Samples
Submitted
Number of QA/QC Samples Submitted
Standards (S) Blanks
(B)
Duplicates
(D)
Pulp Checks 1 SBD Checks 2
2011 10,071 302 310 152 955 60
2012 10,912 339 338 170 1,077 42
2013 16,432 509 514 348 1,827 140
2014 14,826 461 463 404 1,555 153
2015 5,546 173 175 148 549 51
2016 3,923 122 118 83 425 42
2017 5,258 163 163 97 554 62
2018 13,297 432 412 391 1,329 109
2019 4,108 126 126 85 312 20
2020 3 225 8 8 2 0 0
2021 4,563 163 161 98 394 41
Total 129,836 4,035 3,996 2,516 12,835 945

 

  Notes:1 Represent pulps from primary pulp sent to secondary accredited lab
   2 Represent control samples from primary lab sent to secondary lab if enough material was available
   3 In 2020 drilling was only conducted along the MTT alignment, no exploration/resource drilling completed

 

In general, very few control samples assayed between 2006 and 2018 failed (i.e., were outside of +/- 3 standard deviation units of the expected value). Many of the failures were attributed to sample labeling errors (e.g., wrong standard listed in drill log). In the early years, some of the river gravel blanks had anomalous traces of copper and/or gold. The ¼ core duplicate sample results provide some insight into sample reproducibility. Table 11.6 summarizes basic statistics associated with original and ¼ duplicate samples collected at KSM between 2006 and 2021.

 

Table 11.6¼ Core Duplicate Sample Statistics (2006-2021)

 

Parameter Gold (g/t) Copper (%)
Original Duplicate Original Duplicate
Count (@ 0 Cutoff) 4,897 4,897 4,896 4,896
Mean 0.404 0.401 0.126 0.1236
Maximum 9.690 7.970 3.510 3.260
Minimum 0.002 0.002 0 0
Variance 0.251 0.249 0.031 0.031
CV 1.239 1.245 1.403 1.399

 

Figure 4.1 show two box plots (gold on the left and copper on the right) that compare the original ¼ sample with the duplicate ¼ sample.

 

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Figure 11.1 Duplicate Sample Box Plots

 

 

Source: (Wood, 2022)

 

Figure 11.2 and Figure 11.3 show a number of graphical comparisons between the original and duplicate ¼ samples collected at KSM from 2006 to 2021 for gold and copper, respectively.

 

Figure 11.2 Gold Duplicate Sample Graphs

 

 

Source: (Wood, 2022)

 

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Figure 11.3 Copper Duplicate Sample Graphs

 

 

Source: (Wood, 2022)

 

11.3.1Pre-2012 Quality Control Procedures

 

Seabridge developed quality control procedures for their 2006 drilling campaign that remain similar to their current protocols. The quality control procedures included the insertion of certified standard reference materials (SRMs), blanks, and field duplicates into the regular drill core sample stream. During the first two drill campaigns (2006 to 2007), “barren” river gravels collected near Stewart, BC, were used as blank material. In subsequent years, commercially available landscaping materials (crushed marble and granite) were used as blanks. Blanks were inserted into the sample stream at a rate of about one blank for every 33 regular samples. Rare anomalous blank assays for copper and/or gold were attributed to the uncontrolled nature of the blank material.

 

Pre-packaged certified standards, primarily from CDN Resource Laboratories Ltd. (CDN Resource), were submitted at a rate of about one standard per 33 regular core samples. These standards were selected based on the expected gold and copper grades and sample matrix makeup. In 2011, Seabridge had two custom standards prepared and certified by CDN Resource using material from the Mitchell deposit (SEA-KSM) and felsic gold bearing material from a Seabridge deposit located in the Northwest Territories (SEA-CL). Seabridge attempted to ensure that their quality control measures included the insertion of at least one sample blank and one standard within each Eco Tech laboratory batch of approximately 35 samples. The blank and pulp standards were numbered using the same number sequence that was used for the core samples and inserted into each batch shipment randomly by the geologist during the sample layout process.

 

Duplicate field samples consisted of ¼ core samples were collected at a frequency of about one duplicate sample for every 50 regular samples. The HQ and/or NQ core was initially sawn in half and then the original and duplicate sample were created by sawing a specific interval of the half core. The ¼ core pieces were then submitted as original and duplicate samples.

 

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At the conclusion of the drilling/assaying season, Seabridge’s geologic staff selected about 5% to 10% of the primary laboratory assay pulps, which were then sent to an accredited secondary laboratory for “check assay” purposes.

 

Seabridge tried to ensure that they had at least two control samples per the primary laboratory’s fire assay batches, which typically was 33 samples (the laboratory reserved slots for their own control samples). If more than one control standard (SRM or blank) was outside of three standard deviation units of the expected value, the entire batch was deemed suspect and an order was given to the laboratory to re-assay that entire batch. If a single control sample failed and all of the regular samples were unmineralized, then no action was taken. If a single standard failed and the batch contained some mineralized intervals, the control sample plus three to five samples above and below the control sample were re-analyzed. Very few batches were re-run during the period 2006 to 2011.

 

11.3.2Post-2012 Quality Control Procedures

 

Seabridge’s post-2012 quality control procedures were essentially the same as those described for the pre-2012 data and have also been described in previous Technical Reports (Lechner, 2014; Tetra Tech, 2016). Standards, blanks, and field duplicates were submitted at similar frequencies as prior years. Control sample results were monitored by Seabridge personnel during each annual drilling campaign. In most cases where a control sample failed, ten samples on either side of the failed control sample were re-run. In rare cases where more than one standard failed in a batch consecutively, all the samples between the standards were rerun or entire batches were rerun.

 

11.4QP’s Opinion

 

In the opinion of the QP responsible for this section of the Report, sample security, sample preparation, analytical procedures, and QA/QC protocols/results associated with Seabridge’s 2006 to 2021 KSM drilling campaigns were adequate and consistent with standard industry practices. The QP also believes that the assays are suitable to be used to estimate Mineral Resources.

 

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12.0Data Verification

 

12.1Introduction

 

Michael J. Lechner has been the independent QP that performed audits and taken responsibility for the databases supporting the Mineral Resources estimates on the KSM property since April 2007. Mr. Lechner has summarized the data verification procedures he has performed in each of the previous technical reports that are on file on SEDAR. In anticipation of Wood’s QP taking over responsibility for the property description, history, geological setting and mineralization, deposit types, exploration, drilling, sample preparation, analyses and security, data verification, and Mineral Resources, Mr. Lechner did a complete handover of the information on the Kerr, Sulphurets, Mitchell, and Iron Cap deposits to Wood’s Technical Director of Mineral Resource Estimation & Geometallurgy in the 2nd quarter of 2020.

 

12.2Data Verification by Wood Qualified Persons

 

As part of this handover of information, Wood’s Technical Director of Mineral Resources & Geometallurgy visited the KSM site between October 12 and 15, 2019, and August 8 and 12, 2021. Key tasks performed during site visits included:

 

2019 site visit:

 

̶Verification of field geology, logging, and grade data

 

2021 site visit:

 

̶Witness sampling and gold and copper assaying

 

̶Verification of field geology, representative drill collar locations, logging and grade data at Stewart core yard and on the KSM site

 

̶Snowfields check sampling and assaying

 

̶Downhole survey checks on Mitchell-Snowfields dataset

 

̶Quality control checks on Mitchell-Snowfields dataset

 

̶Visual verification of Au, Cu, Ag, Mo, S, As, Zn, Pb data

 

̶Check of logging versus model lithology coding.

 

Additional data verification checks were performed by Wood’s Technical Director of Mineral Resources & Geometallurgy on QA/QC procedures and bulk density measurements.

 

Henry Kim, the current Wood QP responsible for the history, geological setting and mineralization, deposit types, exploration, drilling, sample preparation, analyses and security, and Mineral Resources estimates reviewed the above verification results with Wood’s Technical Director of Mineral Resources & Geometallurgy.

 

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Henry Kim then visited the KSM site and Seabridge’s core storage facility in Stewart, BC, between May 23 and 26, 2021. Key tasks performed during site visits included:

 

Review of drill cores from all five sites, Kerr, Sulphurets, Mitchell, East Mitchell, and Iron Cap, a total 13 representative drill holes

 

Review of core logging, cutting, and sampling procedures, facility, and equipment

 

Visited drill core, pulp reject, and original paper log storage facility in Stewart, BC

 

Reviewed geology and mineralization presentation by Seabridge geology staff

 

Reviewed geology and assay database with Seabridge geology staff

 

Henry Kim was not able to check drill collar locations during his site visit in 2022 due to heavy snow accumulation at KSM site, however, the drill collar location checks by Wood’s Technical Director of Mineral Resources & Geometallurgy are considered adequate.

 

12.3QP’s Opinion

 

12.3.1Drill Hole Data Verification

 

Based on the QP’s review of the extensive data verification procedures carried out and documented by the previous QP on the KSM property, the detailed data verification procedures that were performed by Wood’s Technical Director of Mineral Resources & Geometallurgy (which the current QP reviewed with Wood’s Technical Director), and the current QP’s own site visits and other data verification procedures detailed above, the current QP considers the database adequate to support mineral resource estimation.

 

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13.0Mineral Processing and Metallurgical Testing

 

13.1Introduction

 

The Property includes five major mineralized zones, identified as the Mitchell, East Mitchell (upper zone and main zone), Kerr (upper Kerr and lower Kerr), Sulphurets, and Iron Cap (upper Iron Cap and lower Iron Cap) deposits. The main values in the mineralization of the deposits includes gold, copper, silver, and molybdenum. Principal sulphides in the KSM mineralized zones are pyrite and chalcopyrite. Lower portions of the Kerr and Iron Cap deposits contain minor secondary copper minerals.

 

Extensive metallurgical test programs have been carried out to determine the metallurgical characterization of the deposits (Table 13.1). The test results were summarized in detail in the previous reports by Tetra Tech in 2016 and 2020). Most of the mineralization zones have generally high pyrite-to-chalcopyrite ratios. Pyrite is a significant gold carrier, and copper-to-gold ratios vary significantly in the various KSM deposits. The 2021/2022 metallurgical test work programs focused chiefly on the major mineralized sources of Mitchell and East Mitchell which comprise the majority of the ore sources for this PFS report.

 

Table 13.1Typical Mineralogical Characteristics and Average Copper-to-Gold Grade Ratios in Potential Mill Feed

 

Deposit Average Pyrite/Chalcopyrite
Ratio1
Average
Copper Grade, %
Average
Gold Grade, g/t
Copper/Gold Grade Ratio
Mitchell2 12 0.17 0.61 0.27
East Mitchell – Main5 34 0.13 0.84 0.16
East Mitchell – Upper5 197 0.03 2.03 0.015
Sulphurets2 n/a 0.22 0.59 0.38
Iron Cap3 8 0.35 0.58 0.61
Kerr4 5 0.49 0.31 1.59

Notes:1 Approximate figures.
22016 PFS average mill feed grades.
3Includes Lower Iron Cap.
4Lower Kerr only.
52021 test results only, copper minerals include all copper sulphide minerals.

 

The average Bond Ball Mill Work Index (BBWi) of the Mitchell, East Mitchell and Iron Cap samples range from 13.9 to 15.9 kWh/t which indicates ore that is moderate hardness to ball mill grinding, The Sulphurets mineralization is much harder with an average Bond ball mill work index of 18.5 kWh/t. Table 13.2 shows the average BBWi and abrasion index for the KSM deposits.

 

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Table 13.2Average Ball Mill Grindability and Abrasion Index

 

Deposit Average BBMWi*,
kWh/tonne
Average Bond Abrasion
Index, g
Mitchell 14.4 0.29
East Mitchell 15.5 -
East Mitchell upper 15.7 -
Sulphurets 18.5 0.23
Iron Cap 15.7 0.10
Lower Kerr 14.3 0.06

 

  Note: *Most of tests were conducted at a screen aperture of 106 µm

 

Mineral samples from the Mitchell Zone were tested in the earliest test programs starting in 2007 to develop the process flowsheet for the KSM process design. The flowsheet has been further optimized and then tested on mineral samples from the other zones, in particular the copper and gold recovery by flotation. In 2021, further test work was conducted on samples from Mitchell deposit to verify metallurgical performances of the mineralization from the open pit mine stages which are expected to be mined in the initial mine life years. Also, a separate metallurgical test program was conducted on the samples generated in 2021 from the upper and main zones of East Mitchell deposit. With the verification that the various mineralized samples were amenable to the previous flowsheet, the metallurgical testwork proceeded with the earlier flowsheet and forms the basis of the current KSM processing plant.

 

Table 13.3 summarizes the flotation locked cycle test results from various samples and test programs.

 

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Table 13.3Locked Cycle Flotation Test Result Summary

 

  Mitchell East Mitchell – Main Zone East Mitchell – Upper Zone
Min  Max Average Min  Max Average Min  Max Average
Head Grade
 - Copper, % Cu 0.12 0.24 0.20 0.04 0.22 0.13 - - 0.03
 - Gold, g/t Au 0.55 0.92 0.73 0.48 1.34 0.87 - - 1.34
Cu-Au Concentrate
 - Grade, % Cu 20.2 30.1 24.9 16.8 27.0 24.1 - - 9.3
 - Recovery, % Cu 71.5 89.3 84.5 78.9 91.0 84.9 - - 45.3
 - Recovery, % Au 43.2 73.5 60.8 42.8 67.5 56.0 - - 30.6
  Sulphurets Upper Kerr (Before 2012) Lower Kerr 2013–2017
Min  Max Average Min  Max Average Min  Max Average
Head Grade
 - Copper, % Cu 0.16 0.46 0.26 0.59 0.69 0.62 0.26 1.83 0.64
 - Gold, g/t Au 0.50 0.70 0.59 0.22 0.25 0.24 0.21 0.94 0.46
Cu-Au Concentrate 
 - Grade, % Cu 22.7 29.3 26.9 22.3 30.7 27.8 21.0 28.7 25.2
 - Recovery, % Cu 60.6 85.7 78.0 80.6 86.3 83.0 81.5 96.6 89.7
 - Recovery, % Au 40.1 58.6 52.9 37.7 49.7 41.8 50.5 77.1 62.3
  Iron Cap    
Min  Max Average            
Head Grade
 - Copper, % Cu 0.03 0.59 0.32            
 - Gold, g/t Au 0.20 1.28 0.56            
Cu-Au Concentrate
 - Grade, % Cu 19.2 28.5 24.7            
 - Recovery, % Cu 81.4 93.6 88.5            
 - Recovery, % Au 45.0 74.8 61.9            

 

The cyanide leaching tests of the gold-bearing pyrite tailings (first cleaner scavenger tailings and pyrite concentrate) produced indicated additional gold and silver recovery into a gold doré to be 11% to 17% and 13% to 16% for Mitchell and East Mitchell respectively.

 

Leaching test results suggest that the gold and silver recovery of the gold-bearing sulphide products from the lower Kerr and lower Iron Cap deposits did not seem to respond well to the gold recovery by the established cyanide leaching treatment, compared to the samples from Mitchell, East Mitchell and Sulphurets deposits. Further leaching test work is required to optimize the flowsheet for additional gold and silver recovery, especially from the East Mitchell, Kerr and Iron Cap deposits.

 

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13.2Summary of Metallurgical Test Programs

 

Since 1989, extensive metallurgical test programs have been carried out to determine the metallurgical characterization of the deposits (Table 13.4). Test work from 1989 to 2016 includes all the deposits established and presented in the 2022 PFS. The most recent metallurgical tests were performed in 2021. This recent research assesses and establishes metallurgical recovery methods with fresh samples from Mitchell and East Mitchell deposits. The following sections summarize the test work prior to 2020 with a more detailed review on the most recent test programs.

 

Table 13.4 Metallurgical Test Work Programs

 

Report Year Program ID Lab Mineralogy Flotation/
Cyanide Leach
Grindability Others
2021/2022 KM6486 ALS  
2022 - BQE      
2021 KM6461 ALS
2021 KM6359 ALS    
2020 KM6004 ALS
2019 KM5806 ALS
2018 KM5501 ALS      
2018 KM5501 ALS
2018 KM5455 ALS      
2017 KM5367 ALS    
2017 KM5266 ALS  
2017 KM5248 ALS  
2017 KM5204 ALS      
2017 KM5063 ALS
2016 KM5087 ALS      
2016 SSW30216SD SSW      
2015 KM4514 ALS
2015 KM4672 ALS
2015 - Pocock      
2014 KM4029 ALS
2013 KM3735 ALS    
2013 12628-002 SGS      
2012 KM3174 G&T    
2012 KM3080 G&T    
2011 KM3081 G&T      
2011 KM 2897 G&T      
2011 SSW47110 SSW      
2010/2011 KM 2748 G&T
2010 KM 2755 G&T  
2010 KM 2670 G&T    
2010 #1005610 PRA      
table continues…

 

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Report Year Program ID Lab Mineralogy Flotation/
Cyanide Leach
Grindability Others
2010 #0907510 PRA    
2010 #0904606 PRA    
2009/2010 KM 2535 G&T    
2009/2010 12157-001 SGS    
2009/2010 12248-001 SGS    
2009/2010 #20001153 Metso      
2009/2010 KMA913snR Köeppern      
2009 KM 2344 G&T
2009 - Pocock      
2009 50035-003 SGS      
2009 #0900701 PRA    
2008 KM 2153 G&T
2008 #0607810 PRA      
2008 - Hazen      
2007 KM 1909 G&T
1991 - PDRC  
1990 - PDRC
1989 - BMML    
1989 - CRI      

 

Notes:ALS – ALS Metallurgy Kamloops
 

BMML – Brenda Mines Ltd. Metallurgical Laboratory

BQE – BQE Water

CRI – Coastech Research Inc.

G&T – G&T Metallurgical Services Ltd. (now ALS Metallurgy Pty Ltd.)

Köeppern – Köeppern Machinery Australia Pty Ltd.’s HPGR Pilot Plant at UBC

Metso – Metso Minerals Industries Inc.

PDRC – Placer Dome Research Centre

Pocock – Pocock Industrial Inc.

PRA – Process Research Associates Ltd.

SGS – SGS Mineral Services

SSW– Surface Science Western

 

13.3Summary of Initial Test Work 1989–1991

 

The metallurgical responses of Kerr mineral samples were assessed by initial metallurgical test programs performed by Coastech and Placer Dome between 1989 to 1991. The copper head grade varied from 0.40% to 1.30% Cu. The associated gold and silver concentrations were 0.26 g/t and 1.21 g/t Au and 0.9 g/t to 3 g/t Ag, respectively.

 

The historical test work involved mineralogy, grindability and metallurgical responses to flotation concentration processes. The major conclusions are listed as follows:

 

preliminary mineralogy analysis identified sericite in 3 of the 4 tested samples

 

indicative work index was determined by using a comparative grindability method, which indicates that the tested samples were soft to intermediate hardness

 

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preliminary open circuit flotation tests produced varying metallurgical results with the best results obtained by Placer Dome in 1991 producing salable copper concentrates with significant gold values. However, excessive gold losses to the flotation tailings were reported.

 

13.4Summary of Test Work 2007–2022

 

The 2007 to 2022 metallurgical tests were performed on mineralized samples from Mitchell, Each Mitchell, Sulphurets, upper and lower Kerr, and Iron Cap deposits.

 

13.4.1Test Programs

 

The following test programs were performed to establish a metallurgical flowsheet and to optimize process-related parameters:

 

mineralogy, flotation, cyanidation, and grindability test work by G&T Metallurgical Services Ltd. (G&T) and SGS Minerals Services (SGS). Process Research Associates Ltd. (PRA) also tested some East Mitchell core samples for a previous owner.

 

Metallurgical variability responses and copper-molybdenum separation techniques were also investigated. Flotation locked cycle tests were performed on the composite samples from all the deposits, particularly on a variety of samples from the Mitchell deposit. Cyanidation tests were conducted to further recover gold and silver from the gold-bearing sulphide streams (scavenger cleaner tailing from the copper-gold bulk flotation and an additional pyrite concentrate)

 

semi-autogenous grinding (SAG) mill comminution was evaluated with (SMC) grindability tests to determine the grinding resistance of the mineralization to SAG/ball milling by both Hazen Research Inc. (Hazen) and G&T.

 

crushing resistance parameters for evaluation of high-pressure grinding rolls (HPGR) crushing of the Mitchell and Sulphurets ore samples was also performed by SGS. Pilot scale HPGR testing on Mitchell ore sample was achieved by Köeppern Machinery Australia Pty Ltd.’s (Köeppern) HPGR pilot plant at UBC

 

dewatering tests by Pocock Industrial Inc. (Pocock) and also by PRA were derived on various KSM product pulp samples.

 

13.4.2Baseline Test Process Flowsheet and Conditions

 

The test results indicate that the mineral samples from the five separate mineralized deposits are all amenable to the flotation-cyanidation combined process. The process flowsheet consists of the following stages:

 

copper-gold-molybdenum bulk rougher flotation followed by gold-bearing pyrite flotation

 

regrinding the bulk rougher concentrate followed by 3 stages of cleaner flotation to produce a copper-gold-molybdenum bulk cleaner flotation concentrate

 

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molybdenite separation from the bulk cleaner flotation concentrates to produce a molybdenum concentrate and a copper/gold concentrate containing associated silver. The content of molybdenite is variable and the molybdenum flotation circuit would only operate when the molybdenite content of the bulk concentrate is economical.

 

cyanide leaching of the gold-bearing pyrite flotation concentrate and scavenger cleaner tailing to further recover gold and silver values as doré bullion.

 

The flotation reagents used in the testing were 3418A (dithiophosphinates), A208 (dithiophosphate), and fuel oil for copper-gold-molybdenite bulk flotation, with A208 and potassium amyl xanthate (PAX) used for gold-bearing pyrite flotation. The primary grind size used was 80% passing 125 µm to 150 µm, and concentrate regrind size was 80% passing approximately 20 µm.

 

13.4.3Mitchell Deposit Test Results

 

Mitchell Zone samples were mainly tested by G&T, SGS, Metso, and Hazen from 2007 to 2012 in various metallurgical programs. In 2021, two further metallurgical test programs were conducted to verify the metallurgical performances of Mitchell samples. The KM6359 sample testing program used assay rejects produced from the 2018 drill program and the (KM6461) was generated from the 2021 drill program from the proposed initial mine pits. The KM6359 test results concluded that the fine particle sizes of the assay rejects were likely oxidized due to a lengthy storage time. Further flotation tests were suspended on these “oxidized” samples.

 

MITCHELL SAMPLE CHARACTERISTICS

 

Chemical Composition and Mineralogy

 

The chemical compositions of the Mitchell Zone samples are summarized in Table 13.5. The copper grade is between 0.07% and 0.71% Cu, with an average level of 0.21% Cu; the gold content varies from 0.01 g/t to 1.49 g/t Au, with an average value of 0.79 g/t Au; the silver content varies from 1 g/t to 18 g/t Ag, with an average value of 4 g/t Ag; and the molybdenum content averages approximately 0.007% Mo.

 

The dominate copper mineral identified in the Mitchell samples is chalcopyrite; the main sulphide mineral is pyrite, which was present as approximately 6% to 8% of the sample weight. The degree of chalcopyrite liberation ranged from 46% to 56% across the samples tested at a primary grind size of 80% passing 116 µm to 136 µm. This liberation degree is considered reasonable for rougher flotation.

 

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Table 13.5 Test Samples – Mitchell (2007–2022)

 

Year/Test
Program

Sample
Quantity

Cu (%) Au (g/t) Ag (g/t) Mo (%)
Min Max Avg. Min Max Avg. Min Max Avg. Min Max Avg.
2007/KM1909 3 0.19 0.22 0.20 0.86 0.90 0.87 3 4 4 n/a n/a n/a
2008/KM2153 34 0.07 0.52 0.23 0.22 1.49 0.79 1 18 4 0.001 0.025 0.005
2009/KM2344 10 0.10 0.27 0.22 0.60 1.08 0.93 3 8 4 0.001 0.006 0.007
2009/KM2535 11 0.10 0.71 0.24 0.35 1.02 0.69 2 9 4 0.002 0.015 0.008
2010/KM2670 3 0.12 0.20 0.16 0.01 0.79 0.65 2 3 3 0.006 0.013 0.010
2012/KM3174 25 0.12 0.30 0.20 0.56 1.10 0.76 2 5 3 0.002 0.015 0.005
2021/KM6461 12 0.13 0.31 0.22 0.61 0.96 0.79 3 6 5 0.002 0.003 0.002
Overall 0.07 0.71 0.21 0.01 1.49 0.78 1 18 4 0.001 0.025 0.006

 

Crushability/Grindability – Bond Ball Mill Work Index

 

Over the course of several programs, standard Bond ball mill work index tests were determined on Mitchell composite samples. The results indicate that the Bond work indices range from 12.5 kWh/t to 15.5 kWh/t, averaging 14.4 kWh/t. This suggests that the Mitchell samples are of moderate hardness. The Bond abrasion index (Ai) of Composite PP1 was measured at 0.293 g.

 

Crushability/Grindability – SAG Mill Comminution (SMC) Index

 

SMC grindability tests indicate that the Mitchell samples are moderately resistant to SAG mill grinding with A x b values from 30.0 to 59.6. Based on these results, JK SimMet simulations were conducted for a comminution circuit of 120,000 t/d at an assumed 92% grind circuit availability, and a feed particle size of 80% passing 150 mm. Three scenarios were simulated with different BWi and product size levels as presented in

 

Table 13.6.

 

Table 13.6 JK SimMet Simulation Results (60,000 t/d SABC Circuit, 2008)

 

Simulation 1a 1b 2a 2b 3a 3b
BBWi, kWh/t 14.8 14.8 16.0 16.0 15.0 15.0
SAG Mill Size, D x L (EGL) (ft x ft) 40 x 24 37.7 x 21 40 x 24 37.7 x 21 40 x 24 37.7 x 21
Circulation Load (% of Feed) 19.5 18.4 19.5 18.4 19.5 18.4
Gross Power Draw (kW) 18,843 15,570 18,843 15,570 18,843 15,570
Transfer Particle Size, µm 2,500 3,035 2,500 3,035 2,500 3,035
Ball Mills Size, D x L (EGL) (ft x ft) 22 x 36 22 x 36 22 x 36 22 x 36 22 x 36 24 x 38
Mill Number 2 2 2 2 2 2
Gross Power Draw kW) 15,644 17,293 16,912 18,695 19,283 21,017
Product Size P80, µm 150 150 150 150 120 120
Total Power Draw (kW) 34,487 32,863 35,755 34,265 38,126 36,587
Cyclone Diameter (inches) 26 26 26 26 26 26

 

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Crushability/Grindability – HPGR

 

HPGR laboratory test work was conducted by SGS on both Mitchell and Sulphurets samples. This initial HPGR test work was followed by a pilot plant scale test at Köeppern’s HPGR pilot plant at UBC. The test results indicate the following:

 

the bench scale LABWAL tests by SGS showed that both Mitchell and Sulphurets materials are amenable to HPGR crushing. On average, the net specific energy requirement is 2.33 kWh/t for the Mitchell sample and 3.08 kWh/t for the Sulphurets sample. The Sulphurets mineralization is more resistant to HPGR crushing than the Mitchell mineralization

 

a lower net specific energy consumption (approximately 1.94 kWh/t) was recorded from the closed-circuit pilot plant tests, in comparison with 1.99 kWh/t obtained from the single pass pilot plant testing. A specific pressing force of 4 N/mm2 was considered to be optimum on the basis of both size reduction and throughput performance. An increase in feed moisture is expected to result in a reduction in throughput and an increase in energy consumption.

 

Crushability/Grindability – Tower Mills/Isa Mills

 

Metso investigated the specific energy consumption for secondary ball mill grinding, using tower mills, in jar mill tests. The feed particle size was 80% passing 173 µm, and the product particle size was 125 µm. From the test results, Metso projected the specific energy requirement by a stirred tower mill would be approximately 0.88 kWh/t.

 

SGS used the IsaMill™ procedure to determine the specific energy requirement for regrinding the gold-bearing pyrite rougher concentrate produced from the Mitchell samples. The specific energy requirement to regrind the concentrate from 80% passing 66 µm to 80% passing 16 µm was determined to be approximately 24.2 kWh/t. The grinding media consumption was estimated to be 6 g/kWh.

 

Mitchell Cu-Mo Bulk Concentrate Flotation Test Results

 

Open Cycle Tests

 

In the 2008 testing program, 32 drill core interval composite samples were used for variability tests, excluding two samples (Met 35 and Met 36) from the Sulphurets Zone. Primary grind sizes ranged from 80% passing 115 µm to 171 µm, averaging 149 µm. The rougher concentrates from the copper circuit were reground to approximately 80% passing 18 µm prior to cleaner flotation. The variation in the copper and gold flotation performance of various Mitchell mineral samples is shown in Figure 13.1. The results indicate some variability in copper and gold metallurgical responses, generally related to head grade changes.

 

As shown in Figure 13.2, G&T established the relationship between copper recovery and copper feed grade at a fixed cleaner concentrates grade of 25% copper. This figure indicates that feed grade is the main factor causing copper performance variability.

 

Besides the gold reporting into the final copper concentrate, as shown in Figure 13.1, additional gold recoveries to the gold-bearing pyrite concentrate from the pyrite flotation circuit averaged approximately 16% of the gold in the composite heads. Combined gold recoveries from both the copper-gold flotation circuit and gold-bearing pyrite flotation circuit ranged from 73% to 96%, averaging approximately 86%.

 

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Figure 13.1 Copper and Gold Open Cycle Flotation Variability Test Results (KM2153)

 

 

Source: KM 2153, G & T Metallurgical Services Ltd., Kamloops, BC, Canada, September 16, 2008

 

Figure 13.2 Copper Open Cycle Flotation Performance vs Copper Head Grade at a Concentrate Grade of 25% Copper (KM2153)

 

 

Source: KM 2153, G & T Metallurgical Services Ltd., Kamloops, BC, Canada, September 16, 2008

 

Additional open batch cleaner flotation tests were conducted on nine composite samples representing the major Mitchell Zone mineralization types projected to be mined during various operating periods of an open pit mine plan. The tests were conducted at primary grind sizes, averaging 80% passing 119 µm and regrind sizes of 80% passing 18 µm. The test results show that the average flotation metal recovery was 84.6% for copper and 61.2% for gold from these nine Mitchell metallurgical composites. The test results are shown in Figure 13.3.

 

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The seven composites produced concentrates higher than 25% copper except for 16.2% concentrate copper grade from the low grade IARG 0-10 composite and 24.0% concentrate copper grade from the QSP LG 0-10 composite.

 

Figure 13.3 Flotation Performance – Open Circuit Flotation Tests, Mitchell (KM2153)

 

 

Source: KM 2153, G & T Metallurgical Services Ltd., Kamloops, BC, Canada, September 16, 2008

 

The relationship between the adjusted copper recovery and copper feed grade is plotted in Figure 13.4 with a copper recovery adjustment to reflect a concentrate grade of 25% copper. The graph for the batch flotation tests indicates a good correlation of copper recovery with copper head grade. Although some metallurgical performance variation was observed, the Mitchell mineral samples responded well to conventional flotation.

 

Figure 13.4 Copper Recovery vs. Copper Feed – Open Cleaner Circuit Tests (KM2153)

 

 

Source: KM 2153, G & T Metallurgical Services Ltd., Kamloops, BC, Canada, September 16, 2008

 

Locked Cycle Tests

 

Fifteen locked cycle tests (LCTs) were conducted on the various Mitchell composite samples. The test results are summarized in Table 13.7 for the Mitchell mineralization and in Table 13.8 for Mitchell mineralization samples blended with the other mineralization.

 

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Table 13.7 Locked Cycle Test Results – Mitchell

 

Test
Program*
Comp Grind Size
(P80 µm**)
Feed Grade Bulk Concentrate Grade Flotation Recovery
Cu
(%)

Au

(g/t) 

Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(%)
Ag
(%)
Mo
(%)
G&T 2153/141 Master 119/16 0.21 0.89 4.2 - 20.2 62.8 273 - 87.8 63.0 58.5 -
G&T 2153/142 Master 119/17 0.21 0.92 3.7 - 22.0 64.7 242 - 87.0 58.5 52.5 -
G&T 2344/73 PP Comp 1 103/14 0.24 0.81 - - 22.3 55.7 - - 89.3 66.2 - -
G&T 2535/18 PP Comp 1 103/16         28.0 77.8 260 - 87.2 67.4 47.0 -
G&T 2535/20 PP Comp 1 137/17 0.24 0.82 3.9 - 23.8 62.0 248 - 88.1 66.2 55.6 -
G&T 2670/12 PP Comp 3 147/15 0.20 0.74 3.2 0.006 30.1 77.7 264 0.39 84.2 58.0 52.6 35.7
G&T 2670/18 PP Comp 3 147/22 0.20 0.79 3.2 0.006 27.4 70.5 272 0.46 86.1 56.5 53.0 49.7
G&T 2670/22 PP Hi Mo 143/21 0.16 0.60 3.3 0.014 22.4 61.7 243 1.20 78.9 56.9 43.8 47.9
G&T 2670/26 BS Hi Mo 143/17 0.12 0.55 2.4 0.010 24.9 70.3 185 1.26 71.5 43.2 26.0 42.2
G&T 2897/01 Comp 46 of KM2344 120/16 0.15 0.65 2.3 0.012 22.6 80.5 226 1.76 89.1 73.5 58.6 86.3
G&T 3081/93 Mitchell 3081-M2# 137/18 0.20 0.65 4.7 0.004 27.1 58 427 0.33 83.3 56.3 57.1 55.6
G&T 3081/82 Mitchell 3081-M3# 123/22 0.21 0.57 3.5 0.004 23.8 44.2 223 0.24 88.2 59.9 49.5 49.2
G&T 3081/103 Mitchell 3081-M3# 123/17 0.22 0.55 4.0 0.006 29.8 56 299 0.27 76.7 57.8 43.0 26.3
G&T 6461/16 M-21-152 130/16 0.27 0.77 5.0 0.003 25.0 57.4 251 0.12 87.2 70.8 46.6 45.8
G&T 6461/33 M-21-152 130/15 0.26 0.79 4.0 - 26.1 59.1 280 - 86.6 65.5 59.9 -
G&T 6461/22 Mitchell-2021 HG MC1 126/14 0.26 0.90 4.0 0.002 25.4 68.1 300 0.12 86.0 66.1 59.5 42.3
G&T 6461/23 M-21-151 134/17 0.24 0.91 6.0 0.002 23.3 69.1 444 0.10 89.8 70.4 66.8 45.8
G&T 6461/24 M-21-153 116/15 0.21 0.74 5.0 0.003 22.5 57.0 226 0.23 89.2 63.7 37.6 68.9
G&T 6461/29 Mitchell-2021 HG MC2 127/15 0.19 0.71 4.0 0.002 28.0 75.6 281 0.13 81.2 57.0 41.6 29.5
ALS 4514/30*** Mitchell 4515-M1# 133/15/15 0.21 0.90 5.0 0.006 26.7 98.2 431 0.72 81.6 68.6 53.5 70.0
SGS PP Comp 1 129/28 0.21 0.72 - 0.005 23.1 53.7 - 0.41 89.0 59.6 - 65.0

Notes:*Au grade in the Bulk Cleaner Tailings ranged from 1.01 g/t to 2.26 g/t Au, averaging at 1.63 g/t Au; Au grade in the pyrite concentrate ranged from 0.58 g/t to 2.26 g/t Au, averaging at 1.46 g/t Au.
 **Primary grind size/regrind size
 ***Including a copper flotation on the pyrite flotation concentrate
 #Composite samples IDs based on the previous mine plans (planned in 2010/2011) have been relabeled: Mitchell M1 for Mitchell Yr 0-5. Mitchell M2 for Mitchell Yr 0-10; Mitchell M3 for Mitchell Yr 0-20.

 

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Table 13.8 Locked Cycle Test Results – Blended Samples (Mitchell and Other Deposits)

 

Test
Program*
Comp Grind
Size
(P80
µm**)
Head Grade Bulk Concentrate Grade Flotation Recovery
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(%)
Ag
(%)
Mo
(%)
G&T 2535/19

Mitchell (PP Comp1)/

Kerr (52/53 Blend); 80%:20%

127/20 0.31 0.70 3.5 - 25.3 40.0 168 - 87.4 60.4 51.4 -
G&T 2670/62 Mitchell/Sulphurets Blend; 60%:40% 141/22 0.22 0.67 2.8 0.007 24.2 52.0 178 0.664 85.9 59.8 50.9 72.4
G&T 2748/18

Mitchell (PP Comp 1)/
Iron Cap C1/Iron Cap
C2; 33%:33%:33%

135/15 0.24 0.79 - 0.004 27.6 59.6 - 0.250 87.8 58.2 - 51.5
ALS 4672/32***

Mitchell (Mitchell M1#)/Iron Cap (IC-2014-MC4)

117/17 0.24 0.67 4 0.005 25.0 54.3 304 0.430 82.7 65.7 58.4 70.3
ALS 4514/31*** Mitchell (Mitchell M1#)/ Kerr (DK-2014-MC3) 129/17 0.37 0.59 3 0.006 24.5 28.3 150 0.328 87.9 62.8 57.2 75.1
G&T 3081/81 Mitchell (Year 1-10) /East Mitchell Upper Blend; 90%:20% 110/19 0.17 0.68 3.5 0.008 26.2 61.8 275 0.72 88.8 53.1 46.0 51.1
G&T 3081/94 Mitchell (Year 1-20) /East Mitchell Upper Blend; 90%:20% 125/20 0.16 0.84 3.0 0.005 26.0 83.8 254 0.52 85.2 51.2 43.4 49.9
ALS 6486-59 Mitchell/East Mitchell Upper Blend; 90%:10% 105/15 0.25 0.95 5.5 0.004 27.5 81.0 387 0.230 84.5 65.0 52.7 46.4
ALS 6486-60 Mitchell/East Mitchell Upper Blend; 80%:20% 111/14 0.23 1.01 3.7 0.004 30.1 91.0 302 0.27 85.1 57.9 51.5 40.5
ALS 6486-77 Mitchell/East Mitchell Upper Blend; 70%:30% 106/16 0.18 1.34 4.5 0.006 24.3 111 361 0.58 88.3 55.0 53.1 67.9
Notes:*Au grade in the bulk cleaner tailing ranged from 1.07 g/t to 1.89 g/t Au, averaging at 1.45 g/t Au; Au grade in the pyrite concentrate ranged from to 0.41 g/t to 1.85 g/t Au, averaging at 1.15 g/t Au.
 **Primary grind size/regrind size
 ***Including a copper flotation on the pyrite flotation concentrate
 #Sample ID relabeled, see note in Table 13.7

 

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The LCTs showed that:

 

a substantial variation in the concentrate grade, from 20% to 30% copper, was noticed for individual Mitchell samples. On average, the final copper concentrate contained approximately 25% copper. The average recoveries to the concentrate were 85.1% for copper, 62.1% for gold, 50.7% for silver, and 50.7% for molybdenum

 

approximately 25% of the gold and 26% of the silver in the feed reported to other gold-bearing products, which can be further extracted by cyanide leaching

 

for the blended samples, the metallurgical performance appeared comparable to that produced when treating the Mitchell material on its own.

 

G&T conducted pilot tests on Mitchell samples in 2009 that produced lower metal recoveries and concentrate grades compared with the LCTs results. According to G&T, this was caused by the issues with the pilot plant control or circuit stability. The pilot test results were considered as reference data only and produced bulk samples for other metallurgical process testing.

 

Mitchell Copper-Gold and Molybdenum Separation Flotation Tests

 

Copper-gold and molybdenum separation flotation tests were performed to further recover molybdenum concentrate from copper-gold-molybdenum bulk concentrates in the 2009/2010 testing program. The results indicate that:

 

the best results produced a 51% molybdenum concentrate with a molybdenum recovery of 72% from the bulk copper cleaner concentrate.

 

the molybdenum-copper separation LCT showed 88.5% molybdenum recovery from the molybdenum-copper concentrate at 41% Mo concentrate. The test results are provided in Table 13.9.

 

Table 13.9 Cu-Mo Separation LCT Results, 2010

 

Product Weight
(%)
Grade (%) Recovery (%)
Cu Mo C Cu Mo
Bulk Concentrate 100.0 19.3 1.28 0.63 100.0 100.0
Mo Concentrate 2.8 2.66 41.2 5.76 0.4 88.5
Cu Concentrate 97.2 19.8 0.15 0.48 99.6 11.5

 

Furthermore, G&T conducted preliminary leaching tests on the molybdenum concentrates to reduce the contained copper and lead concentrations. The Brenda-leach method reduced the copper and lead contents from 2.06% to 0.26% for copper and from 0.14% to 0.03% for lead.

 

The assay on the final molybdenum concentrates indicated that the concentrates contained approximately 2,200 g/t rhenium (Re) which is much higher than other molybdenite concentrate producers.

 

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Mitchell Cyanide Leach Tests

 

Cyanide leach tests on the gold-bearing products (first cleaner scavenger tailing and pyrite concentrate) generated from the flotation tests on Mitchell samples were conducted during the time period of 2007 through 2016. Additional tests have been conducted after 2017, on the samples generated during the 2010 pilot plant campaigns, to evaluate cyanidation and other potential treatments. In 2021, further cyanidation tests were conducted to verify the gold and silver extraction from the gold bearing sulphide material generated from the 2021 LCT tests.

 

Mitchell Cyanide Leach Test Samples

 

Cyanide leach tests were conducted on products from the open circuit flotation tests, locked cycle flotation tests, and pilot plant tests. The tests mainly used Carbon-In-Leach (CIL) procedures and Direct Cyanidation (DCN) procedures while the 2021 test work also used a combination of DCN+CIL procedure. The combined first cleaner tailing and the gold-pyrite concentrate were reground and subjected to cyanide leaching for additional gold and silver recovery.

 

For leaching of the open circuit flotation products, the average extraction of gold in these products during e 24-hour leach period in the 2008 program ranged from 63% to 91%, with an average of 79%. For leaching of the LCT products, the average extraction rate is 66% for gold and 59% for silver (Table 13.10).

 

In general, the leach kinetics is rapid for the samples tested. Some of the copper associated with copper minerals was also extracted in cyanidation procedures.

 

Table 13.10 Cyanidation Test Results on LCT Products – Mitchell

 

Testing
Program
Sample Regrind
Size (P80 µm)
Feed
(Au, g/t)
Extraction
(Au, %)
Feed
(Ag, g/t)
Extraction
(Ag, %)
G&T-KM2153 Master 15 1.8 67.6 9.1 62.1
G&T-KM2153 Master 15 2.2 73.2 10.1 64.4
G&T-KM2344 PP Comp 1 12 1.6 68.0    
G&T-KM2535 PP Comp 1 15 1.7 69.0 12.6 54.4
G&T-KM2535 PP Comp 1 15 1.6 81.1 10.9 54.7
G&T-KM2670 PP Comp 3 21 1.6 61.6    
G&T-KM2670 PP Comp 3 18 2.0 66.5 8.1 55.5
G&T-KM2670 PP Hi Mo 19 1.9 68.0 8.6 50.6
G&T-KM2670 BS Hi Mo 19 1.7 68.9 7.6 48.7
G&T-KM2897 Comp 46 - 1.1 63.5 - -
G&T-KM3081 Mitchell M2* 24 1.5 51.2 - -
G&T-KM3081 Mitchell M3* 21 1.2 50.4 - -
SGS PP Comp 1 16 1.1 69.8 - -
G&T-KM6461 Comp M-21-152 (T38) 20 1.3 65.2 9.3 66.5
G&T KM6461 Comp M-21-152 (T40) 20 1.2 68.0 8.7 64.5
G&T KM6461 Comp M-21-152 (T41)** 20 1.2 63.4 8.5 55.2
table continues…

 

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Testing
Program
Sample Regrind
Size (P80 µm)
Feed
(Au, g/t)
Extraction
(Au, %)
Feed
(Ag, g/t)
Extraction
(Ag, %)
G&T KM6461 Comp M-21-152 (T39)** 20 1.4 67.2 9.4 70.2
G&T KM6461 Comp M-21-152 (T42)** 20 1.2 58.8 8.2 65.8
Average – Mitchell 18 1.6 66.2 9.4 59.2
Notes:*Sample ID relabeled, see note in Table 13.7
  **Tests KM6461-39 and KM6461-42 were conducted at a higher cyanide dosage of 2,000 ppm; Tests KM6461-41 and KM6461-42 at a reduced leach retention time of 12 hours and the results are not used for average

 

Some of the leaching tests were conducted separately on the reground first cleaner tailing and reground pyrite concentrate. The overall gold and silver extractions achieved from the separate leaching were similar to the ones from the blend of the two sulphide products.

 

The following major observations are made from the leaching test results:

 

in general, the gold bearing sulphide materials respond reasonably well to cyanidation for recovering residual gold and silver

 

the first cleaner tailing produced lower gold extraction than the pyrite concentrate

 

the gold and silver mineralogical investigation by a Dynamic Secondary Ion Mass Spectrometry (SIMS) technique shows most of the gold that losses after cyanidation occurs as sub-microscopic gold

 

reducing the initial cyanide concentration from 1,000 ppm to 500 ppm generally led to lower extractions of gold, silver, and copper, especially for pyrite concentrate samples

 

longer leaching time was beneficial for silver extraction. Compared with direct leaching process, silver extraction rates were improved significantly by using the CIL test procedure

 

cyanide and lime consumptions for bulk cleaner tailing was significantly higher than pyrite rougher concentrate samples.

 

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13.4.4East Mitchell Deposit Test Results

 

In December of 2020, Seabridge announced the purchase of the Snowfield (SF) deposit which was renamed to East Mitchell, which is immediately adjacent to Mitchell deposit. There are two different primary types of mineralization in East Mitchell deposit:

 

Upper Zone, containing relatively low content of copper minerals and higher values of gold, molybdenum and other base metals, gold values are associated primarily with sulphides including chalcopyrite, pyrite and arsenopyrite

 

Main Zone (also identified as North Zone), showing similar mineralogical characteristics as the Mitchell deposit.

 

Between 2006 and 2011, several test work programs were completed on the samples collected from the East Mitchell deposit. PRA conducted four test programs, labeled as Projects #0607810 (including mineralogical determination studies), #0900701, #0904606, and #0907510. These test programs investigated various potential process routines including direct cyanidation, direct cyanidation with bio-oxidation and pressure oxidation. It also included copper and gold flotation followed by cyanidation of gold bearing flotation by-products. Some tests also incorporated gravity pre-concentration in an effort to recover liberated nugget gold grains. PRA’s test work indicated that East Mitchell’s Main Zone mineralization responded reasonably well to a flotation and cyanidation combined flowsheet, while the low copper grade material from the Upper Zone would require blending with ores containing higher copper grades to enable production of a sellable copper concentrate. PRA’s test work used different flotation test protocols than the currently proposed KSM flowsheet, PRA’s test results are excluded from the review work.

 

In 2011, a separate test work program was completed by Seabridge on various samples generated from East Mitchell deposit. G&T G&T) conducted the test program (KM3801). The test work was focused on determining whether the East Mitchell samples were amenable to the flowsheet established for the other KSM deposits.

 

To compare the metallurgical performances of the East Mitchell samples, two composite samples were constructed from Mitchell deposit, representing the initial 10-year mill feed and 20-year mill feed generated from the upper area of the proposed Mitchell pit based on a preliminary mine plan developed before 2010.

 

Upper Zone Snowfield-Mitchell blended samples were also constructed to investigate the metallurgical responses of these specific blended samples, compared to the metallurgical response of Mitchell Year 0 to 10 sample treated by itself. The blended samples comprised mixes of:

 

10%, 20%, and 50% East Mitchell Upper Zone SFUS and SFUD samples, respectively, with Mitchell (Year 0 to 10 composite)

 

10%, 20%, and 50% East Mitchell Main Zone SFMS, SFMD, and SFULT samples, respectively, with Mitchell (Year 0 to 10 composite).

 

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In 2021, G&T conducted a further test program on samples from the 2021 drilling campaign: These sample composites were prepared from holes SF-21-04, SF-21-05, SF-21-06, and SF-21-08 through SF-21-12.

 

East Mitchell Samples Metallurgical Characteristics

 

Chemical Composition and Mineralogy

 

The chemical compositions of the East Mitchell samples are summarized in Table 13.11.

 

For the Upper Zone samples, the copper grade is between 0.02% and 0.05% Cu with an average of 0.03% Cu; the gold content varies from 1.20 g/t to 2.51 g/t Au, averaging 1.83 g/t Au; the concentrations of silver are relatively constant with an average of 1.5 g/t Ag, and the molybdenum content averages approximately 0.007% Mo.

 

For the Main Zone samples, the copper grade is between 0.05% and 0.19% Cu with an average of 0.13% Cu; the gold content varies from 0.46 g/t to 1.46 g/t Au, averaging 0.80 g/t Au; the concentrations of silver are relatively constant with an average of 2.4 g/t Ag, and the molybdenum content averages approximately 0.006% Mo.

 

Table 13.11Test Samples – East Mitchell (2011 and 2021)

 

Year/Test
Program

Sample

Quantity 

Cu (%) Au (g/t) Ag (g/t) Mo (%)
Min Max Avg Min Max Avg Min Max Avg Min Max Avg
2011 (G&T) - Upper 2 0.03 0.03 0.03 1.42 1.60 1.51 2 2 2 0.012 0.012 0.012
2011 (G&T) - Main 3 0.11 0.14 0.13 0.65 0.75 0.69 2 3 3 0.008 0.009 0.008
2021 (G&T) – Upper 5 0.02 0.05 0.03 1.20 2.51 2.04 1 2 1 0.009 0.022 0.015
2021 (G&T) – Main 8 0.05 0.19 0.12 0.46 1.46 0.76 1 5 3 0.002 0.012 0.008
Overall 0.14 0.46 0.28 0.26 0.81 0.61 1 2 1 0.003 0.011 0.006

 

The copper deportment by QEMSCAN shows higher percentages of copper as copper arsenic sulfosalts for three of the low-grade Upper zone composites. Chalcopyrite was the predominant copper sulphide mineral for the other composites. Estimated copper sulphide liberations at a nominal 80% passing 120 μm by QEMSCAN BMAL protocols were generally lower than those measured in the previous Mitchell studies, Figure 13.5 shows copper sulphides and pyrite liberation status for the East Mitchell samples.

 

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Figure 13.5Copper Sulphides and Pyrite Liberation

 

 

Note: Cus – Copper Sulphides; Gn – Gangues, Py – Pyrite; Ten – Tennantite; En – Enargite

Source: ALS Metallurgy (2021). Metallurgical Testing, KSM Project (KM 6486). July 22, 2022

 

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Crushability/Grindability – Bond Ball Mill Work Index

 

The Bond ball work test results indicate that the East Mitchell mineralization is slightly more resistant to ball mill grinding compared to the Mitchell samples. The average Bond ball work index is 14.8 kWh/t.

 

East Mitchell Cu-Mo Bulk Concentrate Flotation Test Results

 

G&T conducted the open cycle flotation tests using the flowsheet developed for the Mitchell mineralization. No optimization tests were conducted on the East Mitchell samples although some minor modifications were made during the 2021 testing. The test work indicate that on average, the East Mitchell main zone mineralization produced comparable copper metallurgical performances, compared to the Mitchell copper performance, excluding one of the samples from the landslide zone (Composite SF-SNLS-2021-01) producing a low copper grade concentrate. However, the gold flotation performances appear to be slightly inferior, compared to the average Mitchell samples. Also, some significant variations in the bulk flotation performances were observed, especially the samples from the landslide areas of the main zone.

 

The East Mitchell upper zone copper metallurgical performance is different from the Mitchell samples. However, the blends of the Each Mitchell upper zone (up to 30%) and Mitchell did not show detrimental impacts on the Mitchell copper flotation response. Copper concentrate impurity levels attained with the blended sample were higher than noted with the Mitchell concentrate samples, but still within acceptable levels for copper smelters.

 

Gold recovery to the flotation concentrate for the Mitchell and the East Mitchell Upper Zone blended sample appears to be reduced with an increase in the blend ratio from the East Mitchell upper zone mineralization. The inferior gold metallurgical response is possibly due to substantial difference in the mineralogy between the East Mitchell upper zone and the Mitchell materials.

 

Both the 2011 and 2022 test programs conducted LCTs and results are summarized in Table 13.12.

 

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Table 13.12LCT Results – East Mitchell

 

Test Program Composite

Grind Size

(P80 µm*)

Head Grade Bulk Conc. Grade Flotation Recovery (%)
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu Au Ag Mo
G&T 3081/95 50%SFUS + 50%SFUD 117/16 0.03 1.34 2 0.013 9.25 259 213 2.01 45.3 30.6 18.6 24.6
G&T 3081/78 SFMS Composite 117/21 0.13 0.73 2.7 0.009 24.8 70.1 189 0.53 83.6 42.8 31.8 25.3
G&T 3081/79 SFMD Composite 121/19 0.14 0.69 2.2 0.009 25.1 77.1 179 1.09 87.4 54.2 38.9 62.1
G&T 3081/80 SFULT Composite 135/25 0.11 0.90 3.2 0.007 25.5 177 481 1.04 79.6 67.5 52.1 50.6
G&T 6486/53 SF-Main-2021-08 125/18 0.22 1.15 2.3 0.010 25.7 88.8 150 0.57 91.0 59.5 50.9 54.0
G&T 6486/57 SF-Main-2021-09 122/15 0.18 0.76 1.9 0.008 26.7 84.2 180 0.90 90.2 67.4 58.7 71.2
G&T 6486/58 SF-Main-2021-13 121/14 0.04 0.48 0.7 0.010 16.8 106 193 4.15 78.9 43.4 52.8 79.2
G&T 6486/63 SF-Main-2021-MC3 124/15 0.13 0.79 2.1 0.008 25.1 102 175 1.41 86.8 57.7 37.1 75.1
G&T 6486/65 SF-Main-2021-MC4 120/18 0.18 1.07 2.2 0.007 27.0 108 157 0.67 84.5 56.0 39.3 50.4
G&T 6486/89 SF-SNLS-2021-01 131/16 0.11 0.72 1.9 0.007 18.2 84.2 163 0.40 80.2 55.5 41.5 25.6
G&T 6486/70 SF-SNLS-2021-02 123/16 0.17 0.93 3.6 0.012 26.3 92.0 217 0.09 86.5 55.7 34.2 4.4

 

Note: * primary grind size/regrind size

 

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East Mitchell Cyanidation Leach Results

 

The cleaner tailing and the gold-pyrite concentrate from the LCT flotation tests were subjected to cyanide leaching for additional gold and silver recovery. The test results are provided in Table 13.14. In general, the East Mitchell samples produced similar gold and silver extractions, in comparison with the Mitchell samples.

 

The KM6486 test series from 2021 samples also conducted cyanidation tests on the upper zone master composite sample and flotation products separately. The leach test results are summarized in Table 13.13.

 

Table 13.13Direct Cyanidation _ East Mitchell Upper Zone Master Composite

 

Sample Grind Sizing,
P80, µm
Leach Extraction,
%
Residue Grade, g/tonne Reagent Consumption,
- kg/tonne
 
 
Au Ag Cu Au Ag NaCN Lime  
Head 126 59.1 37.0 18.3 0.84 1.3 0.6 0.4  
90 66.5 44.6 18.5 0.74 1.0 0.8 0.4  
62 74.5 45.8 21.7 0.59 1.0 1.1 0.5  
30 83.7 44.5 23.7 0.38 1.1 3.7 0.3  
Combined Sulphide Conc* 30 83.2 56.4 24.3 1.88 4.1 2.6 0.8  
Flotation Tailing* 18 82.7 50.6 30.6 0.06 0.2 0.6 0.2  

 

Note: * Flotation products from Test 29, KM6468

 

The test results indicates that gold extraction improves substantially with an increase in grind size on the East Mitchell upper zone head sample. The silver extraction rate is less sensitive to the grind size change. At a primary grind size of 80% passing 30 µm, the gold and silver extraction rates are 83.7% and 44.5% respectively. A further review on the process method on the upper zone mineralization is required.

 

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Table 13.14Cyanidation Test Results on LCT Products – East Mitchell

 

Test Program Sample Regrind
Size
(P80 µm)
Feed
(Au g/t)
Extraction
(Au %)
Feed
(Ag g/t)
Extraction
(Ag %)
G&T-3081 SFMS-Clnr Tls. 25 2.09 41.6 - -
G&T-3081 SFMS-Pyrite Conc 21 0.62 64.4 - -
G&T-3081 SFMD-Clnr Tls. 19 1.55 49.6 - -
G&T-3081 SFMD-Pyrite Conc 18 1.06 70.8 - -
G&T-3081 SFULT-Clnr Tls. 23 1.63 50.8 - -
G&T-3081 SFULT-Pyrite Conc 19 0.79 64.6 - -
G&T-3081 Mitchell+SFUS/SFUD (80%:20%)-Clnr Tls 21 1.97 50.3 - -
G&T-3081 Mitchell+SFUS/SFUD (80%:20%)-Pyrite Conc 13 1.54 76.0 - -
G&T-3081 Mitchell+SFUS/SFUD (90%:10%)-Clnr Tls 21 1.90 46.0 - -
G&T-3081 Mitchell+SFUS/SFUD (90%:10%)-Pyrite Conc 20 1.38 74.2 - -
G&T-3081 SFUS/SFUD-Clnr Tls. 18 5.23 73.3 - -
G&T-3081 SFUS/SFUD-Pyrite Conc 17 3.07 78.5 - -
ALS-6486 SF-Main-2021-08-Clnr Tls+Pyrite Conc* 19 2.12 43.6 5.2 67.6
ALS-6486 SF-Main-2021-09-Clnr Tls+Pyrite Conc* 16 1.09 49.9 2.7 63.0
ALS-6486 SF-Main-2021-13-Clnr Tls+Pyrite Conc* 17 1.28 59.5 2.8 60.6
ALS-6486 SF-Main-2021-MC3-Clnr Tls+Pyrite Conc* 16 1.47 61.0 3.2 62.9
ALS-6486 SF-Main-2021-MC4-Clnr Tls+Pyrite Conc* 16 1.81 61.9 4.1 60.8
ALS-6486 SF-SFLS-2021-01-Clnr Tls+Pyrite Conc* 14 1.44 61.0 4.9 67.0
ALS-6486 SF-SFLS-2021-02-Clnr Tls+Pyrite Conc* 15 1.27 54.8 4.5 58.0
ALS-6486 Mitchell+East Mitchell Upper Blend (90%:10%)-Clnr Tls+Pyrite Conc* 15 1.36 73.2 6.8 64.6
ALS-6486 Mitchell+East Mitchell Upper Blend (80%:20%)-Clnr Tls+Pyrite Conc* 15 1.88 72.3 7.0 68.7
ALS-6486 Mitchell+East Mitchell Upper Blend (70%:30%)-Clnr Tls+Pyrite Conc* 17 2.25 63.1 6.5 67.5

 

Note: * by DCN+CIL cyanidation procedure

 

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13.4.5Sulphurets Deposit Test Results

 

G&T tested the Sulphurets Zone samples in PFS 2010 and 2011/2012 studies. Three composite samples from crushed drill cores were tested to investigate the metallurgical responses of Sulphurets mineralization.

 

Sulphurets Samples Characteristics

 

Chemical Composition and Mineralogy

 

The chemical compositions of the Sulphurets Zone samples are summarized in Table 13.15. The copper grade is between 0.14% and 0.46% Cu with an average of 0.28% Cu. The gold content varies from 0.26 g/t to 0.81 g/t Au, averaging 0.61 g/t Au. The concentrations of silver are relatively constant with an average of 1 g/t Ag. The molybdenum content averages approximately 0.006% Mo.

 

Table 13.15Test Samples – Sulphurets (2009-2012)

 

Year/Test Program

Sample

Quantity

Cu (%) Au (g/t) Ag (g/t) Mo (%)
Min Max Avg Min Max Avg Min Max Avg Min Max Avg
2009 (G&T) 3 0.14 0.37 0.25 0.26 0.81 0.53 1 2 1 0.003 0.011 0.007
2011/2012(G&T) 2 0.17 0.46 0.32 0.65 0.76 0.70 1 2 1 0.004 0.008 0.006
Overall 0.14 0.46 0.28 0.26 0.81 0.61 1 2 1 0.003 0.011 0.006

 

Crushability/Grindability – Bond Ball Mill Work Index

 

The Bond ball mill work index test results indicate that the Sulphurets samples are more resistant to ball mill grinding compared to the Mitchell samples. The average Bond ball work index is 18.5 kWh/t; the Bond Ai of the overall Sulphurets composite is 0.233 g.

 

Crushability/Grindability – SMC Index

 

The SMC grindability tests indicate that the Sulphurets samples are more resistant to SAG mill grinding than the Mitchell samples. The A x b values of the tested Sulphurets samples were 41.7 and 38.7, respectively.

 

Crushability/Grindability – HPGR

 

SGS conducted bench scale HPGR tests. The test results indicate that the Sulphurets mineralization is more resistant to HPGR crushing than the Mitchell mineralization. On average, the net specific energy requirement determined by bench tests is 3.08 kWh/t for the Sulphurets sample, compared to 2.33 kWh/t for the Mitchell sample.

 

The preliminary HPGR/ball mill circuit simulation results by SGS suggested that the unit power requirement for the HPGR/ball mill circuit would be approximately 14.8 kWh/t for the Sulphurets mineralization, compared to 10.4 kWh/t for the Mitchell mineralization.

 

Sulphurets Cu-Mo Bulk Concentrate Flotation Test Results

 

The open cycle flotation tests by G&T and SGS indicate that the Sulphurets samples show good metallurgical performance and may produce higher grade copper concentrates than the Mitchell samples. Several LCTs have been conducted on the various Sulphurets composite samples and results are summarized in Table 13.16. SGS test recovered 85.7% copper to the bulk concentrate graded at 22.7% Cu, while lower copper recovery but higher copper grade bulk concentrates were produced from the G&T testing program. The testing on Composite 9 produced a much lower copper recovery compared to the other tests, which could be the result of the low head grade.

 

The average gold recovery for both tests was approximately 56%, excluding the lower gold recovery of Composite 9. Silver recovery in the G&T tests averaged 30%. Molybdenum to the bulk concentrate averaged 69%.

 

The locked cycle flotation test results for the metallurgical performances of the Mitchell-Sulphurets blend sample (60% Mitchell and 40% Sulphurets) are presented in Table 13.8 of the Mitchell mineralization section. Similar performance was observed between the blended samples and Mitchell samples.

 

Sulphurets Cyanidation Leach Results

 

The first cleaner tailing and the gold-pyrite concentrate from the flotation LCT tests were subjected to cyanide leaching for additional gold and silver recovery. The test results are provided in Table 13.17. In general, the Sulphurets samples produced lower gold and silver extractions, in comparison with the Mitchell samples. The best gold extraction obtained was 70.5% by SGS using the CIL leach procedure. The direct cyanide leach test appeared to produce inferior results.

 

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Table 13.16LCT Results – Sulphurets

 

Test Program* Composite

Grind Size

(P80 µm**)

Head Grade Bulk Conc. Grade Flotation Recovery (%)
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu Au Ag Mo
SGS Composite 125/20 0.20 0.66 - 0.007 22.7 49.1 - 0.630 85.7 56.1 - 66.6
G&T 2670/44 Master Composite
(Comp49/50/51)
154/16 0.24 0.52 1.6 0.006 28.3 41.8 82.0 0.701 80.5 53.9 34.3 72.2
G&T 2897/22 Master Composite
(Comp49/50/51)
113/- 0.24 0.50 1.5 0.008 28.4 41.6 71.4 0.850 79.4 55.6 31.5 68.5
G&T 3174/8 Composite 8 121/19 0.46 0.70 1 0.008 29.3 31.4 34 0.227 83.6 58.6 31.1 37.7
G&T 3174/9 Composite 9 127/21 0.16 0.59 2 0.004 26.0 63.7 130 0.170 60.6 40.1 21.3 14.1

 

Notes:* Au grade in the Bulk Cleaner Tailings ranged from 1.82 g/t to 3.55 g/t Au, averaging at 2.38 g/t Au; Au grade in the pyrite concentrate ranged from 0.67 g/t to 1.41 g/t Au, averaging at 1.16 g/t Au.

**Primary grind size/regrind size

 

Table 13.17Cyanidation Test Results – Flotation LCT Products, Sulphurets, 2009–2011

 

Test Program Sample Regrind
Size
(P80 µm)
Feed
(Au g/t)
Extraction
(Au %)
Feed
(Ag g/t)
Extraction
(Ag %)
G&T-2670 Master Composite 16 1.7 40.9 3.7 52.4
G&T 2897 Master Composite - 1.5 34.5 3.3 47.9
Composite 8 – 2011/2012 Raewyn CV 25 1.6 41.7    
Composite 9 – 2011/2012 Lower Hazelton 19 2.5 68.3    
SGS (DCN) Composite - 1.6 51.5 - -
SGS (CIL) Composite - 1.3 70.5 - -

 

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13.4.6Upper Kerr Zone Metallurgical Test Results

 

There are two mineralization zones in the Kerr deposit. The lower Kerr Zone mineralization is assumed to be mined by underground block caving, while the upper Kerr Zone material is assumed to be mined by open pit mining.

 

Early test work from 2010 to 2012 completed by G&T were focused on the samples from the surface Kerr Zone.

 

Upper Kerr Samples Characteristics

 

Four composite samples from the upper Kerr were prepared for metallurgical testing from the drill core intervals. It appears that the upper zone materials may have experienced some oxidation and contained more clay minerals.

 

Chemical Composition

 

The assay data of the composites are presented in Table 13.18. The mineralization of each sample can be found in the previous metallurgical reports.

 

Table 13.18Metal Contents of Composites – Upper Kerr, 2010 (G&T)

 

Composite Mineralization Type Cu
(%)
Au
(g/t)
Mo
(%)
Ag
(g/t)
Comp 52 – 2010 Rubble Zone 0.59 0.22 0.004 2.0
Comp 53 – 2010 Quartz Stockwork 0.61 0.17 0.001 1.5
Composite 10 – 2011/2012 CL Quartz Crackle 0.59 0.26 0.001 1.0
Composite 11 – 2011/2012 QSP Quartz Crackle 0.68 0.29 0.001 2.0

 

Crushability/Grindability – Ball Mill

 

The samples from the upper Kerr Zone are more amenable to ball mill grinding when compared to the Mitchell and Sulphurets mineralization. The average Bond ball mill work index is 13.9 kWh/t.

 

Crushability/Grindability – SAG Mill

 

The 2011/2012 testing program revealed that the grindability of the upper Kerr samples to SAG mill grinding is very similar to the samples from the Mitchell deposit. The A x b values of the two tested samples are very close for upper Kerr and Mitchell, 46.1 and 47.0, respectively.

 

Upper Kerr Cu-Mo Bulk Concentrate Flotation Test Results

 

Upper Kerr samples were tested using the same flotation flowsheet developed with Mitchell and Sulphurets samples.

 

The open circuit batch flotation tests showed that the upper Kerr samples produced better concentrate grades than the Mitchell or Sulphurets samples. Copper recovery produced was slightly lower than the Mitchell or Sulphurets samples at the equivalent copper concentration. Gold recovery for the upper Kerr samples was lower because the gold head grades were considerably lower than the samples from the Mitchell and Sulphurets deposits.

 

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The LCT results, as presented in Table 13.19, indicate that the metallurgical performance of the upper Kerr samples was not as good as that achieved with the Mitchell and Sulphurets samples, despite the higher copper head grades at the upper Kerr zone. The inferior metallurgical performance of the upper zone materials may be caused by oxidation and increased clay mineral contents. On average, the upper Kerr samples produced a 27.8% copper concentrate. The copper and gold reporting to the concentrate were 83% and 41%, respectively. Approximately 51% of the gold reported to the gold-bearing pyrite products (first cleaner tailing and gold-bearing pyrite concentrate). The 2011/2012 test program produced better metallurgical performance from the samples tested than what had been achieved previously.

 

The flotation LCT results for the metallurgical performance of the Mitchell-Kerr blend sample (80% Mitchell and 20% upper Kerr) are presented in Table 13.8 of the Mitchell mineralization section. Similar performance was observed between the blended samples and Mitchell samples.

 

Table 13.19LCT Results – Upper Kerr (G&T)

 

Test Program* Comp

Grind Size
(P80 µm**)

Head Grade Bulk Conc. Grade Flotation Recovery
Cu
(%)
Au
(g/t)
Ag
(g/t)
Cu
(%)
Au
(g/t)
Ag
(g/t)
Cu
(%)
Au
(%)
Ag
(%)
G&T 2535/16 Comp 52 119/15 0.59 0.22 1.9 22.3 4.05 33.5 81.6 38.8 37.6
G&T 2535/17 Comp 53 122/14 0.62 0.25 1.4 29.3 5.58 31.8 80.6 37.7 37.9
G&T 3174/10 Composite 10 124/18 0.59 0.24 2 30.7 7.2 49 86.3 49.7 39.8
G&T 3174/11 Composite 11 130/19 0.69 0.24 3 29.0 5.1 77 83.4 41.1 47.4

 

Notes:*Au grade in the Bulk Cleaner Tailings ranged from 0.51 to 0.97 g/t Au, averaging at 0.70 g/t Au; Au grade in the pyrite concentrate ranged from 0.38 to 0.66 g/t Au, averaging at 0.55 g/t Au.
**Primarygrind size/regrind size

 

Cyanidation Leach Results

 

The first cleaner tailing and the gold-pyrite concentrate from the varied flotation circuits were subjected to cyanide leaching for additional gold and silver recovery. G&T conducted cyanidation tests on the products produced from the locked cycle flotation tests; CIL procedure was used for the leaching process. Test results are provided in Table 13.20.

 

On average, gold extraction from both the gold-bearing products was approximately 57%, slightly lower than the results obtained from the Mitchell samples. The average gold feed grade to the cyanide leach circuit was lower in comparison with the cyanide leach feeds of the Mitchell samples. The test results also indicated that the first cleaner tailing produced slightly lower gold and silver recoveries compared to the gold-bearing pyrite concentrate. The average silver extraction was 32%, which was lower than the average extraction of 56% obtained from the Mitchell samples.

 

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Table 13.20Cyanidation Test Results on LCT Products – Upper Kerr (G&T)

 

Test
Program
Sample Regrind Size
(P80 µm)
Feed
(Au g/t)
Extraction
(Au %)
Feed
(Ag g/t)
Extraction
(Ag %)
G&T 2535/G&T 2897 Comp 52 17 1.1 76.0 5.5 45.8
G&T 2535/G&T 2897 Comp 53 15 0.6 59.7 3.2 18.7
G&T 3174 Composite 10 20 0.6 47.2    
G&T 3174 Composite 11 20 0.6 45.6    
Average – Upper Kerr 18 0.7 57.1 4.4 32.3

 

13.4.7Lower Kerr Zone Metallurgical Test Results

 

Five test programs were completed by ALS Metallurgy from 2013 to 2017 with drill core samples from the Lower Kerr Zone. KM3735, KM4029-B, and KM4514 tests were completed between 2013 and 2015. KM5063 and KM5266 in 2017.

 

Lower Kerr Samples Characteristics

 

Composite samples from the lower Kerr Zone were prepared for metallurgical testing from the drill core intervals.

 

Chemical Composition and Mineralogy

 

The metal assays of the individual and master composites samples are summarized in Table 13.21.

 

Table 13.21Test Samples – Lower Kerr (2012–2017)

 

Year/Test
Program

Sample

Quantity

Cu (%) Au (g/t) Ag (g/t) Mo (%)
Min Max Ave Min Max Ave Min Max Ave Min Max Ave
Individual Test Samples
2012/
KM3735
1 0.92 0.32 5 0.004
2013/
KM4029
6 0.41 1.75 0.89 0.27 1.04 0.62 1 4 3 0.004 0.008 0.006
2014/
KM4514
12 0.25 0.86 0.52 0.12 0.76 0.34 1 4 2 0.002 0.009 0.005
2016/
KM5063
17 0.18 0.86 0.51 0.24 0.58 0.40 0.3 4 2 0.001 0.005 0.003
2016/
KM5266
6 0.22 0.67 0.40 0.25 0.50 0.42 2 4 3 0.001 0.003 0.002
Overall 0.18 1.75 0.56 0.12 1.04 0.41 0.3 5 2 0.001 0.009 0.004
table continues…

 

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Year/Test
Program

Sample

Quantity

Cu (%) Au (g/t) Ag (g/t) Mo (%)
Min Max Ave Min Max Ave Min Max Ave Min Max Ave
Master Test Samples
2014/
KM4514
5 0.46 0.54 0.50 0.20 0.44 0.31 1 3 2 0.003 0.006 0.004
2016/
KM5063
12 0.25 0.67 0.47 0.30 0.84 0.40 0.8 3 2 0.002 0.006 0.003
2016/
KM5266
2 0.25 0.51 0.38 0.35 0.43 0.39 3 4 4 0.002 0.003 0.003
Overall 0.25 0.67 0.47 0.20 0.84 0.37 1 4 2 0.002 0.006 0.003

 

 

With individual samples, the copper concentration varied from 0.18% to 1.75% Cu, averaging at 0.56% Cu. The master samples presented less variations of the copper grade: between 0.25% and 0.67% Cu with an average of 0.47% Cu.

 

The mineralogical composition of some of the lower Kerr head samples indicated that the primary copper sulphide mineral was chalcopyrite with a high pyrite content between 5.5% and 10.9%.

 

Grindability

 

The lower Kerr Zone master samples exhibited moderate to average hardness to ball mill grinding process. The average Bond ball mill work index determined by the KM5063 test program is 14.4 kWh/t, ranging from 13.2 to 15.0 kWh/t, as presented in Table 13.22.

 

Table 13.22Bond Ball Mill Work Index Test Results – Lower Kerr, 2017 (ALS KM5063)

 

DK-2015 Master
Samples
Master 215E Master 215W Master 280 Master 775 Average
BBWi 13.2 15.0 14.6 14.6 14.4

 

Mineralogy

 

The KM5063 and KM5266 mineralogical determination results shown in Table 13.23 indicated that the primary copper sulphide mineral is chalcopyrite which is similar to the Mitchell mineralization. Pyrite content is high, ranging between 5.1% and 10.9%. The resulting pyrite to chalcopyrite ratio is between 2.8:1 and 8.5:1.

 

Table 13.23Mineral Composition Data – Lower Kerr 2016/2017 (ALS)

 

Sample Mineral Composition (%)
Chalcopyrite Pyrite Gangue*
KM5063      
DK-2015 Master 215E 1.7 8.9 89.4
DK-2015 Master 215W 1.4 5.5 93.1
DK-2015 Master 280 1.4 10.9 87.7
DK-2015 Master 775 1.7 10.0 88.3
KM5266      
Mineral Intrusive Blend 1.8 5.1 93.1
Wall Rock Sedimentary 0.8 6.8 92.4

 

Note:*Gangue minerals include iron oxides, quartz, feldspars, muscovite, chlorite, carbonates, biotite, apatite, calcium sulphate, titanium minerals, kaolinite, etc.

 

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As presented in Figure 13.6 and Figure 13.7, liberation rates of the copper sulphide minerals has also been estimated with a range from 56% to 66% (KM5063) and 53% to 61% (KM5266), at a primary grind size of 80% passing approximately 130 µm. Most unliberated copper sulphide particles were associated in binary form with non-sulphide gangue minerals.

 

Figure 13.6KM5063 Copper Sulphides Liberation – Mitchell and Lower Kerr 2017 (ALS)

 

 

Source: ALS Metallurgy (2017). Preliminary Metallurgical Testing – Lower Kerr Zones – KSM Project (KM 5063). January 18, 2017.

 

Figure 13.7KM5266 Copper Sulphides Liberation – Lower Kerr 2017 (ALS)

 

 

Source:ALS Metallurgy (2017), Preliminary Metallurgical Testing – Lower Kerr Zones – KSM Project (KM 5266), April 13, 2017.

 

Surface Science Western (SSW) on the leaching residues from Test KM4514-41 found that the residual gold is present in colloidal type sub-microscopic gold, mainly in pyrite, which occurs in coarse and porous types. SSW pointed out that the pre-treatment by pressure or bio-oxidation would be required to release this locked gold.

 

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Lower Kerr Cu-Mo Bulk Concentrate Flotation Test Results

 

Lower Kerr samples were tested using the same flotation flowsheet developed with the Mitchell, Sulphurets, and upper Kerr samples. In addition to batch open circuit cleaner flotation tests, flotation LCTs were completed, with the results shown in Table 13.24. The LCT results show that the lower Kerr mineralization responded well to the tested flowsheet. The copper recovery ranged from 81% to 97% with an average of approximately 90%, and the gold recovery varied from 56% to 77% with an average of approximately 62%. The flotation copper concentrate grades were between 21% and 29%, averaging 25.2%.

 

Table 13.24Flotation LCT Results – Lower Kerr

 

Test Sample Grind Size
(P80 µm^)
Head Grade Bulk Conc. Grade Recovery
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)

Cu

(%)

Au

(%)

Ag

(%)

Mo

(%)

KM4029-22 DK2013-01 118/15/15 0.59 0.36 1.9 0.007 25.3 11.5 36.5 0.243 92.7 69.2 41.0 78.1
KM4029-23 DK2013-02 119/15/11 0.44 0.23 2.0 0.004 24.4 9.0 48.0 0.157 90.6 62.6 40.0 67.4
KM4029-18 DK2013-03 121/14/10 0.50 0.94 4.0 0.007 23.0 32.9 136 0.215 89.7 68.2 65.3 59.4
KM4029-17 DK2013-04 116/16/- 0.86 0.50 2.0 0.003 26.3 9.5 46.0 0.068 91.2 55.8 69.9 58.6
KM4029-16 DK2013-04 116/15/11 0.82 0.50 2.4 0.004 26.3 10.6 48.0 0.071 93.6 61.5 58.3 54.8
KM4029-26* DK2013-05 127/23/10 1.83 0.93 3.3 0.002 28.7 11.2 43.5 0.010 96.6 74.7 80.5 26.4
KM4029-24* DK2013-05 127/30/11 1.81 0.93 3.1 0.002 22.8 9.4 33.8 0.006 96.6 77.1 82.6 29.6
KM4029-25 DK2013-06 128/15/9 1.44 0.67 3.4 0.004 28.4 9.3 50.0 0.039 92.7 65.5 69.2 50.4
KM4514-23 DK-2014-MC3 124/16/15 0.52 0.32 2 0.004 23.5 10.2 55 0.164 91.4 65.9 58.5 76.4
KM4514-24 DK-2014-MC1 121/16/18 0.55 0.42 2 0.003 24.7 14.8 36 0.091 91.3 72.6 48.7 62.8
KM4514-25 DK-2014-MC2 126/16/16 0.51 0.21 3 0.005 26.4 7.3 74 0.210 86.1 56.9 46.3 67.1
KM4514-26 DK-2014-MC4 115/16/16 0.57 0.47 1 0.003 27.4 18.6 34 0.104 89.5 73.8 48.4 67.1
KM4514-27 DK-2014-MC5 128/16/16 0.50 0.24 3 0.006 26.9 8.8 99 0.266 88.5 59.1 63.9 74.7
ALS 4514/31$  Mitchell M1#/DK-2014-MC3 129/17/17 0.37 0.59 3 0.006 24.5 28.3 150 0.328 87.9 62.8 57.2 75.1
KM5063-10 Overall Master# 121/15/14 0.50 0.36 3 0.003 25.7 12.8 81 0.15 84.1 57.4 49.1 70.9
KM5063-11 Master 215E# 115/16/15 0.47 0.33 2 0.005 23.2 12.7 51 0.24 87.1 67.6 51.2 81.4
KM5063-12 Master 215W# 120/14/14 0.51 0.45 3 0.003 25.8 15.6 103 0.091 87.5 59.6 53.3 61.3
KM5063-13 Master 280# 134/16/15 0.48 0.38 2 0.003 21.6 11.3 63 0.093 84.0 56.6 53.0 53.6
KM5063-14 Master 775# 133/16/15 0.54 0.34 2 0.003 24.5 10.6 68 0.13 81.5 56.7 52.6 71.3
KM5063-64 Fresh 775# 125/17/17 0.51 0.26 2 0.004 25.1 7.7 56 0.11 91.5 55.6 52.6 48.6
KM5063-65 Fresh 280# 130/16/19 0.43 0.34 2 0.002 26.3 12.2 79 0.068 90.4 53.1 51.5 44.1
KM5063-66 Fresh 215W# 122/14/17 0.33 0.38 3 0.002 27.4 18.1 166 0.043 90.1 50.5 56.8 23.0
KM5063-67 Fresh 215E# 120/18/19 0.49 0.28 1 0.003 25.9 9.4 38 0.078 93.1 60.3 47.0 42.6
KM5063-78 Fresh SEDS WR# 128/15/20 0.26 0.74 3 0.003 24.7 42.0 144 0.14 88.8 53.4 54.1 39.5
KM5266-09 Mineral Intrusive Blend Master 127/12/14 0.56 0.39 3 0.002 24.8 11.7 73 0.038 89.8 60.9 54.8 38.6
                       

table continues…

 

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Test Sample Grind Size
(P80 µm^)
Head Grade Bulk Conc. Grade Recovery
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)

Cu

(%)

Au

(%)

Ag

(%)

Mo

(%)

KM5266-10 Wall Rock Sedimentary Master 134/14/15 0.27 0.48 3 0.002 21.0 27.7 192 0.105 85.0 61.6 61.9 46.0

 

Notes:*: Au grade in the Bulk Cleaner Tailings ranged from 0.29 to 1.50 g/t Au, averaging at 0.64 g/t Au; Au grade in the pyrite concentrate ranged from 0.07 to 0.66 g/t Au, averaging at 0.27 g/t Au.
^: Grind Size includes primary grind size/bulk concentrate regrind size/pyrite concentrate regrind size.
$: Repeat tests, the cleaner flotation for Test 26 was conducted on more diluted slurry (using a larger flotation cell)
#: DK2015 Samples

 

Cyanidation Leach Results

 

ALS Metallurgy conducted cyanidation tests on the first cleaner tailing and the gold-bearing pyrite concentrate produced from the flotation LCTs. Test results produced before 2016 are presented in Table 13.25 and after 2016 in Table 13.26.

 

Table 13.25Preliminary Cyanidation Test Results – Lower Kerr (before 2016)

 

Test Sample Test Procedure* Extraction
Gold (%) Silver (%)
KM4029-29 Test 22,23 Bulk Cleaner Scavenger Tailing CIL 72.6 86.2
KM4029-30 Test 22,23 Gold-Bearing Pyrite Concentrate CIL 77.2 90.8
KM4029-31 Test 16,18 Bulk Cleaner Scavenger Tailing CIL 59.1 76.5
KM4029-32 Test 16,18 Gold-Bearing Pyrite Concentrate CIL 73.9 91.2
KM4029-33 Test 24,25 Bulk Cleaner Scavenger Tailing CIL 57.5 73.1
KM4029-34 Test 24,25 Gold-Bearing Pyrite Concentrate CIL 72.2 n/a
KM4514-33 Bulk Cleaner Scavenger Tailing CIL 64.2 82.6
KM4514-34 Gold-Bearing Pyrite Concentrate CIL 77.8 70.8
KM4514-35 Bulk Cleaner Scavenger Tailing CIL 68.1 56.9
KM4514-36 Gold-Bearing Pyrite Concentrate CIL 72.9 78.9
KM4514-37 Bulk Cleaner Scavenger Tailing CIL 66.3 78.2
KM4514-38 Gold-Bearing Pyrite Concentrate CIL 68.4 54.3
KM4514-39 Bulk Cleaner Scavenger Tailing CIL 58.2 75.7
KM4514-40 Gold-Bearing Pyrite Concentrate CIL 63.7 72.1
KM4514-41 Bulk Cleaner Scavenger Tailing** CIL 56.6 71.9
KM4514-42 Gold-Bearing Pyrite Concentrate** CIL 67.1 71.8

 

Notes:* Cyanide concentration: 1,000 ppm; pH: 11; carbon addition: 28 g/L
** Mitchell (Mitchell Year 0-5)/ Kerr (DK-2014-MC3) The mineralogical study by Iron Cap Zone Major Metallurgical Test Results

 

The gold extractions from KM4029 and KM4514 by cyanide leaching (CIL procedure) fluctuated from 57% to 73% for the bulk cleaner scavenger tailing and from 64% to 78% for the gold-bearing pyrite concentrates.

 

Table 13.26Preliminary Cyanidation Test Results – Lower Kerr

 

Test Head Samples Test Procedure* Extraction (%)
Au Ag
Bulk Cleaner Scavenger Tailing Cyanidation
KM5063-15 DK-2015-Overall Master CIL 62.2 83.6
KM5063-16 DK-2015-Master 215E CIL 65.7 71.7
KM5063-17 DK-2015-Master 215W CIL 61.0 55.5
KM5063-18 DK-2015-Master 280 CIL 52.9 77.5
KM5063-19 DK-2015-Master 775 CIL 54.6 82.2
KM5063-69 DK-2015-Seds-WR CIL 58.1 74.8
KM5063-70 DK-2015-775 Comp 2 CIL 61.5 69.6
    table continues…

 

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Test Head Samples Test Procedure* Extraction (%)
Au Ag
KM5063-71 DK-2015-280 Comp 2 CIL 48.2 72.7
KM5063-72 DK-2015-215W Comp 2 CIL 57.1 79.8
KM5063-73 DK-2015-215E Comp 2 CIL 60.8 73.5
KM5063-25 DK-2015-Overall Master DCN 70.7 55.5
KM5063-40** DK-2015-Overall Master DCN 64.2 42.6
KM5063-41** DK-2015-Overall Master DCN 65.8 43.3
KM5063-74 DK-2015-775 Comp 2 DCN 68.1 52.8
KM5063-75 DK-2015-280 Comp 2 DCN 52.3 67.8
KM5063-76 DK-2015-215W Comp 2 DCN 53.2 68.4
KM5063-77 DK-2015-215E Comp 2 DCN 63.9 46.5
KM5063-79 DK-2015-Seds-WR DCN 67.4 82.7
KM5266-11 Mineral Intrusive Blend Master DCN 70.3 93.4
KM5266-12 Wall Rock Sedimentary Master DCN 72.9 78.9
Copper First Cleaner Tailing Cyanidation (Au-Pyrite Concentrate)
KM5063-20 DK-2015-Overall Master CIL 83.5 71.8
KM5063-21 DK-2015-Master 215E CIL 91.5 58.3
KM5063-22 DK-2015-Master 215W CIL 70.2 73.1
KM5063-23 DK-2015-Master 280 CIL 71.9 58.2
KM5063-24 DK-2015-Master 775 CIL 82.3 66.6
KM5266-13 Mineral Intrusive Blend Master DCN 87.9 94.3
KM5266-14 Wall Rock Sedimentary Master DCN 71.2 96.5

 

Notes:*Cyanide concentration: 1,000 ppm; pH: 11; carbon addition: 28 g/L
**Cyanidation was performed on a finer feed particle size

 

Further test work by the KM5063 and KM5266 test programs were conducted on the bulk cleaner scavenger tailing and copper first cleaner tailing (pyrite concentrate) produced from the LCT tests. Both CIL (for KM5063 program only) and DCN leach procedures were used for the tests. The average gold extraction rates from both the tailing samples are higher than previous results.

 

For the bulk cleaner scavenger tailing containing approximately 0.6 g/t to 1.6 g/t gold and 1.8 g/t to 6.3 g/t silver, the extractions fluctuated widely from 48% to 73% for gold, averaging 61.5%, and from 42% to 93% for silver, averaging 68.6%.

 

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For the gold-bearing pyrite concentrates grading at 0.17 g/t to 0.49 g/t gold and 0.9 g/t to 2.0 g/t silver, the extractions also varied significantly, ranging from 70% to 92% for gold, averaging 79.8%, and from 58% to 97% for silver, averaging 74.1%. The gold and silver extraction varies substantially and did not exhibit correlations with the leach feed grades including gold and copper, as well as the regrind size which was maintained at 80% passing 14-15 µm.

 

The CIL test results also show that a significant amount of copper was leached out together with the precious metals. It appears that the copper dissolution had a significant impact on the gold adsorption onto the activated carbon, but no obvious effect on the silver adsorption. On average, only approximately 66% of the extracted gold was captured onto the carbon.

 

In 2016, ALS Metallurgy also completed a metallurgical test program (KM5087) to generate materials for gold and silver extraction testing. The samples generated were the first cleaner scavenger tailings and pyrite concentrates from a blended master composite made up of both the Mitchell and lower Kerr zone samples. ALS Metallurgy conducted preliminary cyanide leach tests after the samples had been treated by various pre-aeration methods. On average, approximately 64% of the gold was extracted. The average silver extraction was approximately 90%, while the copper extraction averaged at approximately 29%.

 

13.4.8Iron Cap Deposit Metallurgical Test Results

 

Previous test programs on Iron Cap were conducted in 2010 and between 2014 to 2015. The 2010 test work was based on two composite samples, while further tests in 2014 and 2015 were performed using samples from lower Iron Cap zone. Four additional metallurgical test programs, KM5248, KM5501, KM5806, and KM6004, were performed by ALS Metallurgy on samples from lower Iron Cap zone through 2017 to 2020.

 

Iron Cap Samples Characteristics

 

Chemical Composition and Mineralogy

 

The assays of the head samples from the Iron Cap zone are summarized in Table 13.27. The average copper and gold grades of the individual Iron Cap samples were 0.29% Cu and 0.54 g/t Au, which is similar with the master sample results. Preliminary mineralogy work was conducted in 2010 and 2015. The results from the two programs indicated that chalcopyrite is the major copper-bearing mineral with a portion of chalcocite, covellite, and tennantite/tetrahedrite. Pyrite is the dominant sulphide mineral. Feldspars, micas, and quartz are the main gangue minerals.

 

Table 13.27Test Samples – Iron Cap

 

Year/Test
Program

Sample

Quantity

Cu (%) Au (g/t) Ag (g/t) Mo (%)
Min Max Ave Min Max Ave Min Max Ave Min Max Ave
KM2748 3 0.14 0.36 0.25 0.32 1.06 0.71 5 6 6 0.002 0.003 0.003
KM4029 3 0.22 0.25 0.24 0.28 0.59 0.45 3 4 4 0.003 0.008 0.006
KM4672 15 0.16 0.47 0.30 0.11 1.72 0.56 1 10 4 0.001 0.007 0.004
KM4672* 4 0.16 0.34 0.27 0.35 0.63 0.49 2 5 4 0.002 0.004 0.003
KM5248 6 0.18 0.86 0.41 0.42 1.30 0.87 3 6 4.5 0.001 0.007 0.003
KM5501 8 0.01 0.72 0.37 0.17 2.91 0.87 1 5 3 0.002 0.016 0.005
KM5501* 3 0.31 0.58 0.44 0.53 0.62 0.57 1 2 1.3 0.003 0.008 0.005
KM5806 6 0.18 0.49 0.35 0.19 0.74 0.49 2 4 2.5 0.002 0.007 0.004
KM6004 12 0.20 0.55 0.36 0.28 1.03 0.55 1 4 2.3 0.001 0.005 0.003
Overall 0.01 0.86 0.30 0.11 2.91 0.55 1 10 3.5 0.001 0.016 0.004

 

·Note: * Composite samples

 

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Grindability

 

Bond ball mill work index tests were conducted in various test programs on the samples from Iron Cap deposit. The overall results are presented in Table 13.28, which shows that the work index ranges from 13.4 kWh/t to 17.4 kWh/t, averaging 15.8 kWh/t. The measured index indicates a moderate hardness to ball mill grinding for the tested Iron Cap samples. The measured abrasion index fluctuates between 0.094 g and 0.200 g, with an average level of 0.147 g.

 

Table 13.28Bond Ball Mill Work Index Test Results – Iron Cap (ALS)

 

Samples Wi
(kWh/t)
Samples Wi
(kWh/t)
Samples

Abrasion Index, Ai

(g)

IC 2010 Composite 1 14.9 IC-2017-MC-B 15.9 IC-2014-MC4 0.099
IC 2010 Composite 2 14.9 IC-2018-01 17.4 IC-2018-13 0.094
IC-2014-MC4 16.5 IC-2018-02 15.8 IC-2018-14 0.130
IC-2017-01 15.9 IC-2018-03 15.6 IC-2018-15 0.143
IC-2017-02 16.1 IC-2018-04 15.9 IC-2018-16 0.186
IC-2017-05 15.6 IC-2018-05 15.6 IC-2018-17 0.200
IC-2017-07 16.6 IC-2018-06 13.4 IC-2018-18 0.175
IC-2014-MC1 16.3 IC-2018-13 15.2    
IC-2014-MC2 15.9 IC-2018-14 15.8    
IC-2014-MC3 16.7 IC-2018-15 15.9    
IC-2016-01/02 17.4 IC-2018-16 15.6    
IC-2016-03/04 16.0 IC-2018-17 15.1    
IC-2017-MC-A 16.0 IC-2018-18 15.7    
Average Wi, kWh/t   15.8 Average Ai, g 0.147
             

The KM4672 program also tested the grindability of the IC-2014-MC4 master composite sample to SAG mill grinding using the SMC procedure. The results show that the SMC parameters are A = 68.7, b = 0.54, and A x b = 37.1.

 

Mineralogy

 

The mineralogical composition and liberations of the composite samples have been included in all the testing programs. Table 13.29 shows the major results. As with the Mitchell and lower Kerr zone samples, the primary copper sulphide mineral identified is chalcopyrite. However, tennantite and enargite have been identified as secondary dominate copper minerals in the sample IC-2018-06.

 

The pyrite content is between 2.1 and 9.2. The pyrite-to-chalcopyrite ratio fluctuates between 1.8:1 and 23:1. Liberation of the copper sulphide minerals at the specified grind size were reported in a range of 53% to 65% (KM5248), 46% to 56% (KM5501), 40% to 52% (KM5806), and 45% to 51% (KM6004). Liberation of less than 50%, based on ALS Metallurgy, is not sufficient to efficiently recover copper minerals at the rougher flotation stage.

 

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Table 13.29Mineral Composition Data – Iron Cap 2017-2020 (ALS)

 

Sample Mineral Composition (%) Sizing Liberation (%)
Chalcopyrite Tennantite/
Enargite
Other Cu
Minerals*
Pyrite Gangue (µm)
P80
Copper
Sulphide
KM5248              
IC-2016-01 2.4 0.1 <0.01 4.4 93.1 116 65
IC-2016-02 1.0 <0.1 <0.01 5.5 93.5 127 53
IC-2016-03 0.6 <0.1 <0.1 2.6 96.8 121 54
KM5501              
IC-2017-08 <0.1 <0.1 <0.1 4.6 95.4 80 56
IC-MC-A 0.7 <0.1 <0.1 3.8 95.5 123 46
IC-MC-B 1.5 0.1 <0.1 4.6 93.8 111 52
IC-2014-MC1 0.9 <0.1 <0.1 4.6 94.5 n/a n/a
IC-2014-MC2 0.6 <0.1 0.0 4.0 95.4 n/a n/a
IC-2014-MC3 0.4 - <0.1 3.0 96.6 n/a n/a
IC-2016-04 0.5 - <0.1 3.0 96.5 n/a n/a
KM5806              
IC-2018-01 0.4 0.0 0.0 3.8 95.8 82 41
IC-2018-02 0.8 <0.1 <0.1 3.3 95.9 95 40
IC-2018-03 1.2 <0.1 0.0 3.6 95.2 86 46
IC-2018-04 0.8 - 0.0 2.2 97.0 85 52
IC-2018-05 1.2 <0.1 0.0 2.1 96.7 85 44
IC-2018-06 0.4 0.5 <0.1 9.2 89.9 80 51
KM6604              
IC-2018-13 1.3 0.2 <0.1 6.2 92.3 114 51
IC-2018-14 1.3 <0.1 <0.1 3.9 94.8 128 47
IC-2018-15 1.1 <0.1 <0.1 3.2 95.7 124 50
IC-2018-16 0.9 <0.1 <0.1 2.1 97.0 124 50
IC-2018-17 1.0 <0.1 <0.1 4.1 94.9 128 45
IC-2018-18 1.0 <0.1 <0.1 4.6 94.4 122 45

 

Note:*Other copper minerals refer to bornite and chalcocite/covellite.

 

Iron Cap Cu-Mo Bulk Concentrate Flotation Test Results

 

Using the same metallurgical flowsheets used for the other deposits, Iron Cap samples were tested with batch open circuit flotation test procedure, including cleaner flotation tests, and locked cycle flotation test procedure.

 

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The primary grind size test results indicate that a slightly finer primary grind size would produce better copper and gold recoveries. The process condition tests also indicates that a lower rougher flotation pulp density (approximately 22% w/w) would improve the overall copper and gold metallurgical performance for some samples, compared to a “standard” slurry pulp density of approximately 35% to 40% w/w. The lower rougher flotation slurry densities resulted in a 30% to 40% reduction in rougher flotation weight.

 

The averaged silver recovery to the flotation concentrate was slightly higher than the recovery achieved from the Mitchell mineralization. On average, approximately 61% of the molybdenum from the Iron Cap samples reported to the final bulk concentrate. As shown in Table 13.8, the Mitchell and Iron Cap blended samples did not show detrimental effects of the blending on the metallurgical responses.

 

Major observations of the results from the Iron Cap LCTs (Table 13.30) are as follows:

 

the tested head samples contain varied contents of copper, gold, and silver, which range between 0.03% and 0.59% Cu, 0.20 g/t and 1.28 g/t Au, 1 g/t and 6 g/t Ag, and 0.002% to 0.007%, respectively

 

on average, approximately 88.5% copper has been recovered to the bulk concentrate product grading at 24.7% Cu. Approximately, 62% gold and 56% silver have been recovered into the flotation concentrate

 

the bulk cleaner scavenger tailing still contain significant gold and silver contents which, on average, represent 24% gold at a grade of 1.1 g/t Au and 21% silver at a grade of 5 g/t Ag

 

on average, the gold-pyrite concentrate contains approximately 4.5% gold at 0.45 g/t Au and 3.7% silver at 2 g/t Ag, respectively.

 

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Table 13.30Flotation Locked Cycle Test Results – Iron Cap 2010 -2020

 

Test* Sample Grind Size
(P80 µm**)
Head Grade Bulk Conc. Grade Recovery
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(%)
Ag
(%)
Mo
(%)
G&T 2748/11 IC-2010 Comp 1 150/15 0.14 1.28 6 0.002 25.4 147 774 0.180 81.6 55.2 61.0 37.9
G&T 2748/12 IC-2010 Comp 2 147/22 0.38 0.31 5 0.003 24.9 10 255 0.115 88.1 45.0 62.0 55.2
G&T 2748/17 50%Comp 1:
50%Comp 2
108/19 0.26 0.82 - 0.003 26.7 51.9 - 0.144 85.2 53.3 - 41.5
ALS 4029/19 IC-2013-01 117/16/7 0.25 0.56 3 0.006 26.7 50.4 179 0.481 83.4 71.4 48.4 69.0
ALS 4029/20 IC-2013-02 119/17/11 0.26 0.51 4 0.004 22.9 36.4 273 0.238 85.4 68.4 62.8 64.4
ALS 4029/21 IC-2013-03 130/15/11 0.23 0.24 4 0.003 23.4 14.8 258 0.205 87.5 53.4 56.3 56.4
ALS 4514/30 IC-2014-MC1 125/14/14 0.33 0.69 4 0.004 25.4 41.5 230 0.238 87.1 67.8 60.1 70.3
ALS 4514/29 IC-2014-MC2 127/14/16 0.25 0.45 4 0.004 23.4 29.3 275 0.311 85.7 59.1 56.0 76.2
ALS 4514/31 IC-2014-MC3 124/16/18 0.16 0.32 2 0.002 22.6 37.6 139 0.187 81.4 67.9 47.4 62.9
ALS 4514/25 IC-2014-MC4 124/14/15 0.28 0.56 4 0.003 24.9 36.2 250 0.257 85.7 62.9 65.1 73.6
KM5248-10 IC-2016-02 87/13/13 0.37 0.45 6 0.006 24.0 20.9 245 0.300 86.9 63.1 59.8 66.6
KM5248-11 IC-2016-03/04 88/13/12 0.20 1.03 3 0.002 22.2 85.5 264 0.113 82.0 61.6 61.1 54.6
KM5501-16 IC-MC-A 123/19/18 0.32 0.68 3 0.002 22.9 34.4 151 0.094 88.3 62.3 67.3 56.5
KM5501-17 IC-MC-B 111/17/16 0.59 0.58 4 0.007 26.8 18.9 137 0.193 92.2 66.8 77.1 54.3
KM5501-19 IC-MC-C 116/16/17 0.43 0.64 3 0.004 27.0 28.5 156 0.132 90.0 64.2 68.2 43.6
KM5806-18 IC-2018-01 100/22/19 0.18 0.20 1 0.006 21.6 13.9 86 0.590 88.4 51.8 42.4 74.2
KM5806-19 IC-2018-02 95/16/14 0.35 0.33 2 0.003 23.1 14.0 60 0.125 90.8 59.1 52.4 60.5
KM5806-20 IC-2018-03 108/20/17 0.42 0.69 2 0.003 23.9 28.3 67 0.105 92.0 67.1 59.7 63.5
KM5806-21 IC-2018-04 100/15/14 0.27 0.35 1 0.005 25.5 26.5 67 0.386 92.1 74.8 50.1 83.2
KM5806-22 IC-2018-05 108/19/21 0.50 0.75 2 0.002 25.2 27.0 64 0.074 92.4 65.7 48.7 55.5
KM5806-23 IC-2018-02 95/17/16 0.35 0.31 2 0.003 23.9 13.2 66 0.135 90.3 56.3 51.9 60.0
KM5806-24 IC-2018-03 108/19/22 0.43 0.69 2 0.002 25.1 28.8 68 0.107 90.7 64.7 61.5 76.2
KM6004-21 IC-2018-13 90/16/16 0.46 0.46 5 0.005 22.7 11.4 150 0.19 92.7 46.8 57.1 70.2
KM6004-22 IC-2018-14 93/19/16 0.40 0.63 3 0.003 19.2 21.7 72 0.12 91.0 65.5 47.2 69.8
KM6004-23 IC-2018-15 94/17/16 0.32 0.51 2 0.003 24.4 26.4 75 0.16 93.6 64.0 58.2 63.6
                        table continues…

 

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Test* Sample Grind Size
(P80 µm**)
Head Grade Bulk Conc. Grade Recovery
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(g/t)
Ag
(g/t)
Mo
(%)
Cu
(%)
Au
(%)
Ag
(%)
Mo
(%)
KM6004-34 IC-2018-14 93/18/20 0.40 0.64 2 0.004 24.3 27.6 84 0.14 90.3 63.4 52.6 57.8
KM6004-35 IC-2018-16 89/15/16 0.027 0.70 2 0.002 26.2 52.5 116 0.03 90.5 70.6 50.9 20.0
KM6004-44 IC-2018-17 86/18/20 0.36 0.45 2 0.002 28.4 23.0 90 0.070 91.9 59.2 52.4 37.8
KM6004-45 IC-2018-18 89/16/17 0.35 0.53 2 0.003 27.6 26.1 109 0.127 91.3 57.4 56.5 59.0
KM6004-58 IC-2018-14 152/16/16 0.40 0.60 3 0.004 27.6 29.7 102 0.16 86.7 63.0 50.8 50.4
KM6004-59 IC-2018-15 152/16/16 0.32 0.46 2 0.003 28.5 30.8 85 0.12 88.0 65.9 47.5 37.4

 

Notes:* Au grade in the Bulk Cleaner Tailings ranged from 0.53 to 2.17 g/t Au; Au grade in the pyrite concentrate ranged from 0.05 to 1.88 g/t Au.
**Primary grind size/regrind size.

 

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Cyanidation Leach Results

 

G&T conducted cyanidation tests on the first cleaner tailing and the gold-bearing pyrite concentrate produced from the flotation LCTs. Test results before 2016 are provided in Table 13.31.

 

Table 13.31Cyanidation Test Results on LCT Products – Iron Cap

 

Testing
Program
Sample Regrind
P80 (µm)
Feed
(Au g/t)
Extraction
(Au %)
Feed
(Ag g/t)
Extraction
(Ag %)
G&T-2748 Iron Cap Comp1 14 1.9 49.7 9.4 62.8
G&T-2748 Iron Cap Comp2 15 1.1 40.4 6.9 56.8
G&T-2748 50% Comp1/50% Comp2 16 1.5 48.6 - -
ALS-4029 IC-2013-01/02/03 Cl.Sc.Tls 16 0.8 45.8 4.4 70.7
ALS-4029 IC-2013-01/02/03 Py Conc 10 0.2 54.7 1.6 87.4
ALS-4672 IC-2014-MC1 Cl.Sc.Tls 14 1.1 46.7 6 68.0
ALS-4672 IC-2014-MC2 Cl.Sc.Tls 14 1.2 29.2 7 69.7
ALS-4672 IC-2014-MC4 Cl.Sc.Tls 14 1.1 40.1 6 60.5
ALS-4672 IC-2014-MC1 Py Conc 14 0.4 50.6 3 57.2
ALS-4672 IC-2014-MC2 Py Conc 16 0.3 36.3 2 72.0
ALS-4672 IC-2014-MC4 Py Conc 15 0.1 74.9 5 73.2
Average – Iron Cap 15 1.0 46.8 5.6 66.5
           

On average, the gold extraction from both the gold-bearing products was approximately 47%. As compared with the Mitchell samples, Iron Cap samples produced lower gold leach extractions, especially for leaching of the first cleaner tailing. The average silver extraction of 67% is higher than the average extraction of 56% obtained the Mitchell samples.

 

Further cyanidation leach tests including DCN leaching and CIL tests have been conducted by ALS Metallurgy during 2017 and 2020. The bulk cleaner scavenger tailing and copper first cleaner tailings produced from the locked cycle testing programs of KM5248, KM5501, KM5806, and KM6004 have been tested.

 

All the cyanide leach test results are provided in Table 13.32. The following conclusions could be made based on the leach test results:

 

·gold extractions from the bulk cleaner scavenger tailing were 35% to 70% with an average of 51%. On average, about 1.9 kg/t of sodium cyanide and 3.4 kg/t of lime were consumed in the leaching process

 

·gold extractions to the leach feed from the copper first cleaner tailing were 40% to 73% with an average of 53%. On average, about 1.6 kg/t of sodium cyanide and 2.6 kg/t of lime were consumed

 

·the average silver extraction rates were approximately 65% for the bulk cleaner scavenger tailing and 66% for the copper first cleaner tailing

 

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·for the composite samples constructed from both the tailing products, gold extraction rates were slightly lower, ranging between 40 and 44%

 

·the CIL tests show that, on average, less than 80% of the gold extracted was adsorbed onto the activated carbon, ranging from 42% to 98%. The low gold adsorption onto the activated carbon might be due to significant copper dissolution in the leach treatment, which might have been caused the inferior gold adsorption onto the activated carbon.

 

Table 13.32Preliminary Cyanidation Test Results – Iron Cap (ALS 2017 - 2020)

 

Test Test Type

Cyanide

Consumption

(g/t)

Lime

Consumption

(g/t)

Feed Grade, Calculated Extraction
Au
(g/t)
Ag
(g/t)
Cu
(%)
Au (%)

Ag

(%)

Cu
(%)
Bulk Cleaner Scavenger Tailing Cyanidation
KM5248-12 DCN 2,800 3,900 0.91 11.45 0.17 70.3 57.2 55.8
KM5248-13 DCN 2,600 3,700 1.99 8.50 0.18 50.8 49.4 54.6
KM5501-21 CIL 1,900 3,000 1.43 4.78 0.07 51.9 60.3 21.9
KM5501-22 CIL 2,700 4,000 1.30 4.06 0.06 51.7 58.1 40.5
KM5501-23 CIL 2,300 4,100 1.45 4.22 0.08 57.9 66.8 37.3
KM5806-18 DCN 1,100 5,400 1.2 4.0 0.05 60.8 64.7 29.8
KM5806-19 DCN 900 4,900 1.1 3.5 0.06 48.6 62.9 25.2
KM5806-20 DCN 800 2,800 1.2 3.3 0.06 43.2 63.8 26.5
KM5806-21 DCN 1,100 3,100 0.9 2.3 0.06 63.3 65.7 34.3
KM5806-22 DCN 900 3,700 1.4 3.7 0.07 43.5 70.1 29.0
KM6004-36 CIL 2,000 3,600 1.17 5 0.07 34.6 57.0 40.1
KM6004-37 CIL 1,300 2,400 1.07 3 0.08 40.0 68.5 20.1
KM6004-39 CIL 2,500 2,700 1.29 4 0.11 45.1 73.2 39.5
KM6004-38 CIL 1,500 2,400 0.76 2 0.05 43.4 74.3 32.1
KM6004-52 CIL 2,500 2,300 1.21 3 0.08 54.5 65.7 43.2
KM6004-53 CIL 2,600 3,100 1.09 5 0.09 58.2 76.8 43.4
KM6004-54 CIL 3,100 3,200 1.03 4 0.09 42.0 70.6 49.1
Copper First Cleaner Tailing Cyanidation (Au-Pyrite Concentrate)
KM5248-14 DCN 1,800 2,000 0.29 2.60 0.10 72.5 73.1 64.6
KM5248-15 DCN 1,700 2,200 0.73 3.28 0.08 53.2 60.3 68.3
KM5501-24 CIL 1,600 1,200 0.46 1.43 0.03 56.9 65.1 22.4
KM5501-25 CIL 2,300 1,300 0.48 1.18 0.05 57.0 74.6 45.3
KM5501-26 CIL 2,000 1,300 0.52 1.64 0.05 54.2 69.4 43.8
KM5806-18 DCN 1,200 4,800 0.87 3.0 0.04 63.8 63.6 33.6
KM5806-19 DCN 1,000 4,400 0.77 2.0 0.04 51.9 60.0 36.7
KM5806-20 DCN 1,000 4,600 0.84 2.1 0.04 49.9 62.5 39.9
KM5806-21 DCN 800 4,400 0.42 1.0 0.03 60.5 71.2 43.5
KM5806-22 DCN 1,100 3,700 0.82 2.3 0.04 40.0 60.7 37.6
KM6004-40 CIL 1,000 2,200 0.26 2 0.03 65.8 70.0 37.5
KM6004-41 CIL 900 2,200 0.32 2 0.02 51.3 44.4 38.5

table continues…

 

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Test Test Type

Cyanide

Consumption

(g/t)

Lime

Consumption

(g/t)

Feed Grade, Calculated Extraction
Au
(g/t)
Ag
(g/t)
Cu
(%)
Au (%)

Ag

(%)

Cu
(%)
KM6004-43 CIL 1,700 2,600 0.50 2 0.03 56.9 68.0 69.5
KM6004-42 CIL 1,100 2,200 0.38 2 0.04 45.7 75.3 30.6
KM6004-55 CIL 2,400 1,800 0.46 1 0.04 54.7 71.5 53.2
KM6004-56 CIL 2,100 2,100 0.54 2 0.05 54.4 66.4 40.9
KM6004-57 CIL 2,200 2,500 0.55 2 0.04 42.0 78.2 52.0
Combined Tailing Cyanidation (Selected Bulk Cleaner Scavenger Tailing+ Copper First Cleaner Tailing)
KM5501-32 DCN 1,200 2,300 0.98 4.0 0.06 40.0 62.7 37.4
KM6004-67 CIL 2,600 2,000 0.74 3 0.07 43.6 58.4 42.0
KM6004-68 CIL 1,900 2,000 0.72 3 0.07 40.4 56.4 39.8

 

13.4.9Flotation Concentrate Assay (2007–2022)

 

The multi-element assay data are presented in Table 13.33 and Table 13.34 for the concentrates from various deposits. On average, the impurities in the copper-gold concentrates produced from the Mitchell, East Mitchell, Sulphurets, Iron Cap, and Kerr deposits should not incur smelting penalties as set out by most smelters, excluding the concentrate from the upper East Mitchell zone.

 

However, arsenic, antimony, and mercury contents in some of the concentrates from the upper East Mitchell, the Iron Cap and the Kerr materials may incur smelting penalties, especially the East Mitchell Upper zone materials. According to the mine plan, the East Mitchell Upper materials will be blended the Mitchell materials. The blending will be controlled and the contribution of the East Mitchell Upper materials to the mill feed is expected to be small. Therefore, it is expected that penalty if any should be minimum. Also, the lead content of the concentrate from the Iron Cap Composite 1 may be higher than the penalty thresholds. Fluorine levels in some of the concentrates may be also higher than the penalty thresholds. It is anticipated that the mill feeds will be supplied from different deposits and mineralization zones. Impurity contents in the copper concentrates produced from these blended mill feeds should be lower than the penalty thresholds set by most of the smelters. Further review with respect to smelting penalties should be conducted.

 

ALS Environmental has conducted additional selenium testing on Iron Cap samples for selenium deportment and distributions on the flotation products, as well as for the effects of cyanidation of selenium extraction. The selenium analysis results are summarized as follows:

 

·about 21% selenium in the sample reported to the bulk concentrate at a grade of 114 g/t Se; very little selenium (<1 ppm) was dissolved into the process water of the rougher and cleaner tailings

 

·a low percentage (about 3%) of selenium in the leach feed was extracted during the cyanidation leach. During the destruction of the cyanide, the residual selenium concentration was 0.186 mg/L as compared with 0.221 g/t Se in the non-treated cyanide leaching residue. The selenium content can be effectively removed by using reverse osmosis (RO) methods.

 

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Table 13.33Flotation Concentrate Assay from Different Deposit Samples

 

Element Unit  Mitchell East Mitchell - Main East Mitchell Upper Sulphurets Kerr Iron Cap
    Min Max Avg Min Max Avg Min Min Max Avg Min Max Avg Min Max Avg
Copper (Cu) % 22.0 28.0 25.0 18.2 27.0 24.9 9.3 26.0 29.3 27.9 21.0 30.7 25.6 21.7 27.0 24.4
Gold (Au) g/t 44.2 98.2 65.1 70.1 177.0 98.2 259 31.4 63.7 45.6 4.1 32.9 12.1 10.9 146.8 35.9
Silver (Ag) g/t 223 431 303 150 481 210 213 34 130 82 32 192 69 59 774 180
Molybdenum (Mo) % 0.12 0.72 0.35 0.09 1.41 0.7 2.01 0.17 0.70 0.37 0.01 0.27 0.11 0.07 0.59 0.20
Zinc (Zn) % 0.23 0.43 0.33 0.2 1.8 0.6 2.46 0.18 0.54 0.34 0.01 0.94 0.30 0.02 2.67 0.78
Arsenic (As) ppm 690 2,080 1,181 200 3,258 1,604 12,020 205 1,768 732 143 3,276 1,575 91 6,140 1,716
Selenium (Se) ppm 59 102 75 95 123 111 158 118 118 118 109 247 205 108 400 255
Antimony (Sb) ppm 210 1,182 647 <100 1,800 530 5,100 370 2,100 972 24 2,710 801.8 22 4,500 1,653
Mercury (Hg) ppm 0.6 6.8 2.8 0.3 5.9 1.4 <1 2.0 2.0 2.0 3.4 88.0 25.5 <1 25.0 6.2
Lead (Pb) % 0.12 0.92 0.30 0.24 0.47 0.36 0.53 0.19 0.72 0.39 0.02 1.33 0.12 0.06 1.31 0.43
Bismuth (Bi) ppm <10 150.0 76.0 25.0 29.0 27.0 90 <10 <10 <10 3.1 105.0 21.5 5.9 205.0 47.4
Fluoride (F) ppm 69 346 168 84 170 132 647 155 155 155 20 264 179 162 494 347
Chloride (Cl) % <0.01 <0.01 <0.01 0.02 0.02 0.02 - <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 <0.01 <0.01

 

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Table 13.34Flotation Concentrate Assay from Blended Head Samples

 

Element Unit Mitchell/
Upper
East
Mitchell
Mitchell/
Sulphurets
Mitchell/
Upper
Kerr
Mitchell/
Lower
Kerr

Mitchell/Iron Cap,
Avg

Copper (Cu) % 26.8 24.2 25.3 24.5 25.2
Gold (Au) g/t 85.7 52.0 40.0 28.3 47.2
Silver (Ag) g/t 316 178 168 150 250
Molybdenum (Mo) % 0.46 0.66 0.06 0.33 0.26
Zinc (Zn) % 0.33 0.92 0.42 0.24 0.6
Arsenic (As) ppm 2,512 969 1,369 1,404 1,670
Selenium (Se) ppm 79 89 76 - 76
Antimony (Sb) ppm 1,200 500 492 814 1,304
Mercury (Hg) ppm 3.9 1.0 2.4 23.0 4.7
Lead (Pb) % 0.2 0.26 0.15 0.12 0.3
Bismuth (Bi) ppm 39.0 <10 121.0 12.2 80.4
Fluoride (F) ppm 134.0 174 116 264 191
Chloride (Cl) % - <0.01 - <0.01 -

 

13.4.10Ancillary Tests

 

During test programs, various environment-related tests have been conducted to determine engineering-related parameters. The key tests are as follows:

 

·leach residue cyanide destruction, including sulphur dioxide/air, Caro’s acid (H2SO5), and hydrogen peroxide (H2O2)

 

·cyanide recovery from barren solutions, including acidification, volatilization of hydrogen cyanide gas, and re-neutralization (AVR); and sulphidization, acidification, recycling, and thickening of precipitate (SART)

 

·static and dynamic thickening tests for conventional thickener sizing and for high rate thickener sizing for primary grinding product, first cleaner tailing together with gold-bearing pyrite concentrate, cyanidation residues, and rougher/scavenger flotation tailing

 

·filtration testing, including vacuum filtration and pressure filtration for bulk flotation concentrate.

 

Cyanide Recovery Tests & Cyanide Destruction Tests – 2009/2010 (SGS)

 

A large-scale, agitated bulk cyanide leach test was conducted by SGS on a 20 kg combined sample of first cleaner tailing and pyrite rougher concentrate. The sample was sourced from material generated from the flotation pilot plant testing at G&T. The key chemical analysis of the solution for cyanide recovery and the washed leach pulp for the cyanide destruction are shown in Table 13.35.

 

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Table 13.35Chemical Analysis of Cyanide Recovery Test Solution and Cyanide Destruction Pulp

 

Sample CNT
(mg/L)
CNWAD
(mg/L)
CNF
(mg/L)
Cu
(mg/L)
Fe
(mg/L)
CNS
(mg/L)
Leach Solution 853 850 280 562 1.6 700
Washed Pulp 94 90 - 90.4 1.08 220

Note: CNT—total cyanide; CNWAD—weak acid dissociable cyanide; CNF—free cyanide; CNS—thiocyanate

 

Exploratory AVR tests were conducted to investigate the effect of pH on the recovery of cyanide from the barren leach solution. The scrubbing retention time was 4 hours. The collected cyanide, acid consumption, and lime consumption are summarized in Table 13.36.

 

Table 13.36Cyanide Recovery Test Results – AVR

 

pH Recovered
CNWAD (%)
Sulphuric Acid
Addition (g/L)
Hydrated Lime
Addition (g/L)
2 77 3.18 0.78
3 72 2.01 0.24
4 35 1.14 0.16

Exploratory SART tests were also conducted on the barren leach solution to investigate the effects of pH and sodium hydrosulphide (NaHS) dosage on recovering cyanide and copper from CNWAD and copper cyanide complexes. The test results are as follows:

 

·at a sodium hydrosulphide dosage of 100% stoichiometric requirement, 83% to 94% of the copper was precipitated when reducing the pH level from 5 to 3

 

·at pH 3, an increase of sodium hydrosulphide dosage to 120% of the stoichiometric requirement resulted in near complete removal of copper from the solution and regeneration of all the weak acid dissociable cyanide as free cyanide

 

·the sulphuric acid addition was approximately 1.9 g/L of feed solution, and the hydrated lime requirement for re-neutralization of the SART-treated solution was 1.3 g/L of feed solution.

 

Further optimization of the SART conditions could improve these results, if SART was considered for recovery of cyanide into low-concentration cyanide solutions. These SART-generated cyanide solutions might also be considered for feed to further AVR processing to generate higher grade cyanide solutions for recycle to the leaching circuits.

 

Cyanide Destruction Tests – 2009/2010 (SGS)

 

Three different cyanide destruction methods — sulphur dioxide/air, Caro’s acid, and hydrogen peroxide — were tested for oxidation of cyanide and detoxification of the washed pulp. The objective of the test work was to produce treated effluent containing less than 2 mg/L CNWAD. The results of the cyanide destruction test results are summarized in Table 13.37.

 

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The results indicated that the residual CNWAD in the washed pulp was reduced to less than 1 mg/L after the pulp had been treated with 4 g to 5 g equivalent sulphur dioxide and 0.14 g copper (added as copper sulphate) per gram of CNWAD in the pulp. The reaction time for this process was one hour at the natural pH. The sulphur dioxide/air-treated pulp contained small amounts of CNT in the form of ferrocyanide complex.

 

An exploratory test indicated that the residual CNWAD in the solution phase of the washed pulp was reduced to less than 2 mg/L level by using Caro’s acid treatment. The reagent consumption was 0.74 g/L H2SO5 (250% of the stoichiometric amount) and 0.6 g/L hydrated lime of the feed to the cyanide destruction.

 

The tests also indicated that the hydrogen peroxide process is not very efficient for cyanide destruction. The residue CNWAD was only reduced from 90 mg/L to 11 mg/L after adding 500% of the stoichiometrically required hydrogen peroxide.

 

Two-stage cyanide destruction involving sulphur dioxide/air treatment followed by a polishing treatment with Caro’s acid or hydrogen peroxide was investigated on the pulp and also on a tailing filtrate solution. The sulphur dioxide/air-treated pulp was adjusted with sodium cyanide to 10 mg/L CNWAD for the polishing tests. The results are as follows:

 

·the polishing test using Caro’s acid was unsuccessful. The final product still contained 3.2 mg/CNWAD after adding 500% of the stoichiometric Caro’s acid

 

·the hydrogen peroxide polishing treatment produced less than 2 mg/L residual CNWAD. The hydrogen peroxide dosage was 10 times of the stoichiometric requirement and the copper addition was 0.011 g/L pulp

 

the solution phase (filtrate) of the sulphur dioxide/air partially treated pulp responded well to the hydrogen peroxide polishing treatment. The solution contained less than 1 mg/L residual CNWAD after being treated with five times the stoichiometric hydrogen peroxide requirement (0.065 g/L solution). Copper sulphate was not used in the treatment of this solution.

 

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Table 13.37Cyanide Destruction Test Results – 2009/2010 (SGS)

 

Test Method Oxidant
Dosage
Stoich (%)
Cumulative
Retention
Time (~h)
Composition (Solution Phase) Cumulative Reagent Addition* (g/g CNWAD)
pH CNT
mg/L
CNWAD
mg/L
SO2
Equivalent
Lime Cu H2SO5
100%
H2O2
100%
Cu mg/L
Solution
Cyanidation Washed Pulp 10.7 94 90 - - - - - -
CND 6&7 SO2/Air 160-200 1 9.6 2-4 <1 4-5 - 0.14 - - 12
C-1 Caro’s Acid 500 1.5** 9.0 2.8 1.7 - 37 - 21.9 - -
H-1 H2O2 500 1.5** 10.1 12 11 - - - - 6.5 -
SO2/Air Partially Treated Pulp 10.0 10 10 -   - - - -
C-2 Caro’s Acid 500 1.5** 9.0 2.8 1.7 - - - 21.6 - -
H-7 H2O2 1,000 0.5 10.0 2.3 0.3 - - 1.5 - 13 15
SO2/Air Partially Treated Solution 10.0 10 10 - - - - - -
H-4 H2O2 500 1 8.7 1.6 0.4 - - - - 6.5 -
Notes:*Copper was added as CuSO4 5H2O; SO2 was added as Na2S2O5

**Reagent was added in three 30-min stages

 

Cyanide Recovery Tests & Cyanide Destruction Tests – 2017 and 2022 (BQE Water)

 

Six leaching residue samples from the test program KM5367 on the Mitchell zone have been tested for cyanide destruction and copper removal using sulphur dioxide/air method. The copper removal was completed after the cyanide destruction to achieve the effluent copper concentration target of 0.5 mg/L.

 

The results confirmed that the cyanide destruction was successful by reducing the WAD cyanide concentration in the effluent to a level between 0.005 mg/L and 0.075 mg/L, as compared with the target limit of 0.5 ppm CNWAD.

 

In 2022, BQE Water (BQE) conducted a preliminary test work to investigate cyanide and copper recovery and gold adsorption onto the carbon. Below are the observations from the testing:

 

·Near complete (~ 99%) removal of copper, and corresponding recovery of free cyanide, was achieved by SART treatment. The reagent consumption was close to stoichiometric values to metals;

 

·Zinc and silver present in the SART feed were removed from solution and reported to the Cu2S solids;

 

·No loss or deportment of gold into solids was found in SART treatment;

 

·The gold recovery from SART effluent in a single pass flow-through activated carbon in column was as high as 92% under both tested alkaline and acidic conditions. This corresponds to reduction in gold concentration from 0.26 mg/L in feed solution to less than 0.03 mg/L in column effluent (barren solution).

 

Settling Tests

 

Preliminary settling tests were conducted on pyrite rougher flotation tailing in the 2008 testing program. As reported by G&T, the tests on the tailing slurry failed to generate normal settling curves. The tests were subsequently carried out on the re-pulped sample from dried tailing. The test data reveal that the settling area required for pyrite rougher flotation tailing was 0.73 m2/t/d without adding flocculant and 0.30 m2/t/d with the addition of 10 g/t of flocculant.

 

In 2009, Pocock conducted solids liquid separation (SLS) tests on five flotation products generated by G&T from the bench scale and pilot plant tests. The materials tested included flotation feed, copper concentrate, first cleaner tails and gold-bearing pyrite concentrate, cyanidation residues, and rougher/scavenger flotation tailing. The dewatering tests included:

 

·flocculant screening tests

 

·static and dynamic thickening tests for conventional thickener sizing and for high rate thickener sizing

 

·viscosity (rheological properties) tests for rake mechanism and underflow pipeline sizing

 

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·vacuum and pressure filtration tests

 

Hychem AF 303 (a medium to high molecular weight, 7% charge density, anionic polyacrylamide) was selected for thickening tests from preliminary screening of a series of flocculants. The key test results are summarized in Table 13.38 and Table 13.39.

 

Table 13.38 Recommended Conventional Thickener Operating Parameters – 2009 (Pocock)

 

Tested Material Feed
(% Solids)
Flocculant
(g/t)
Underflow
(% Solids)
Unit Area
(m2/t/d)
Flotation Feed Composite 20-25 10-15 60-65 0.125
Coarse Grind Flotation Feed 25-30 10-15 70-74 0.125
Final Copper Concentrate 25-30 5-10 70-72 0.125
Rougher Tailing 15-20 10-15 60-62 0.125
Au-Pyrite Concentrate and Cu Cleaner Tailing 15-20 20-25 55-58 0.275-0.307
Cyanide Leach Reside 10-15 20-25 50-53 0.284-0.312
Notes:–All tests were performed at 20°C and as received pH.
–Hydraulic loading or rise rate (m3/m2/h) includes a 0.5 scale-up factor.
–Unit area includes a 1.25 scale-up factor; the range of unit areas provided corresponds to the range of underflow densities.
–Coarse grind flotation feed: at a particle size of P80 170 µm; simulating stage one primary grind size.

 

Table 13.39Recommended High Rate Thickener Operating Parameters – 2009 (Pocock)

 

Tested Material Feed
(% Solids)
Flocculant
(g/t)
Underflow
(% Solids)
Net Feed Loading
(m3/m2h)
Flotation Feed Composite 15-20 15-20 60-65 4.8-6.1
Coarse Grind Flotation Feed 20-25 10-15 70-74 4.8-6.1
Rougher Flotation Tailing 15-20 ~20 57-62 3.7-4.8

 

Filtration Tests

 

The 2009 Pocock testing program also determined the filtration rates of the copper concentrates produced from G&T pilot plant tests. Both vacuum filtration and pressure filtration methods were tested. The test results are summarized in Table 13.40.

 

Table 13.40Filtration Test Results – 2009 (Pocock)

 

Filtration
Method
Bulk Cake Density
(dry kg/m3)
Cake
Thickness
(mm)
Cake
Moisture
(%)
Filtration
Rate
(dry kg/m2h)
Dry Cake
Weight
(dry kg/m2)
Vacuum 1,785 15 19 265* -
Pressure 2,511 51 8 - 117.8**
Notes:*Includes scale-up factors at vacuum of 67.7 kPa.
**Feed pressure 552 kPa at 51 mm thickness.

 

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Magnetic Separation Tests

 

In the 2008 test program, Davis Tube magnetic separation was used in an effort to recover the metal contents lost in the coarser than 74 µm fraction of the pyrite flotation tailing from Tests 10, 11, and 25. Test results indicate that less than 3% of the coarse tailing weight was recovered into a magnetic fraction assaying approximately 23% iron. No copper or gold assay data was reported.

 

13.5Conclusions

 

Extensive metallurgical testing programs were conducted on the samples representing the four deposits of the KSM Property in 1989-1991 and 2007-2022.

 

The substantial test results indicate that the mineral samples from the five separate deposits are amenable to conventional flotation. In general, the Mitchell and East Mitchell, Sulphurets, and parts of the Kerr and Iron Cap mineralization responds well to additional gold and silver recovery from the gold-bearing sulphide products by cyanidation. However, the test results show that the gold-bearing sulphide products (first cleaner scavenger tailing and pyrite concentrate) from the lower Kerr and lower Iron Cap zones did not seem to respond well to gold recovery by using the tested cyanide leaching treatment. Also the mineralization from the East Mitchell upper zone shows different metallurgical responses although the material responds reasonably well to the established KSM process flowsheet, especially with moderate ore blending ratios with Mitchell ores. Further test work is required to optimize the process conditions, or alternative treatment methods should be used for the gold and silver recovery in the East Mitchell upper ore zones.

 

The flotation and cyanidation combined process consists of:

 

·copper-gold-molybdenum bulk rougher flotation followed by gold-bearing pyrite flotation

 

·regrinding of the resulting bulk rougher concentrate followed by three stages of cleaner flotation to produce a copper-gold-molybdenum bulk cleaner flotation concentrate

 

·molybdenum separation of the bulk cleaner flotation concentrate to produce a molybdenum concentrate and a copper/gold concentrate containing associated silver, if the molybdenum grade is sufficiently high

 

·cyanide leaching of the gold-bearing pyrite flotation concentrate and the scavenger cleaner tailing to further recover gold and silver values as doré, mainly for Mitchell, East Mitchell and Sulphurets mineralization and parts of Iron Cap and Kerr mineralization.

 

The testing programs from 2017 to 2022 show the following:

 

·a finer primary grind size can improve the copper and gold metallurgical performance, especially for copper

 

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·rougher flotation at a low slurry solid density can improve copper and gold metallurgical performance, especially for the mineralization with more clay-type minerals from lower Iron Cap and lower Kerr zones

 

·the test results suggest that the gold-bearing sulphide products (first cleaner scavenger tailing and pyrite concentrate) from the lower Kerr and lower Iron Cap zones did not seem to respond well to the gold recovery by the established cyanide leaching treatment

 

·Some of the samples from the East Mitchell deposit, especially the landslide zone and the major KSM fault zones, do not respond well to the established KSM reagent regime. Further tests are needed to optimize process conditions for these “altered” fault zones.

 

·The samples from the upper zone of East Mitchell deposit shows different mineralogical property and metallurgical performances, compared to the other mineralization. Further test work is required to optimize the process conditions and flowsheet for the material.

 

On average, the impurities in the copper-gold concentrates produced from the Mitchell, main East Mitchell, Sulphurets, and Kerr deposits should not incur smelting penalties as set out by most smelters. However, arsenic, antimony, and mercury contents in some of the concentrates from the Iron Cap deposit, the Kerr samples and the East Mitchell Upper samples may incur smelting penalties. It is anticipated that the mill will be supplied with blended feeds from different deposits. Impurity contents in the copper concentrates produced from these blended mill feeds should be lower than the penalty thresholds set by most of the smelters. The blend ratio for the East Mitchell Upper zones should be controlled properly to avoid elevated impurity levels in the flotation concentrate. Further review with respect to smelting penalties should be conducted.

 

The samples from all the deposits are moderately hard for ball mill and SAG mill grinding, excluding the samples from the Sulphurets deposit, which show high energy requirements for both ball mill and SAG mill grinding. All the samples tested are amenable to particle size reduction by HPGR procedure.

 

13.6Metallurgical Performance Projection

 

The metallurgical test results obtained from the various test programs were used to predict the plant metallurgical performance parameters for copper, gold, silver, and molybdenum. Gold and silver recoveries are based on the combined process of flotation to produce a salable concentrate, followed by cyanidation of combined cleaner tailings and pyrite flotation concentrate. The flotation process will produce a copper concentrate containing approximately 25% copper with variable precious metal contents and a molybdenum concentrate with 50% molybdenum. The gold cyanidation process applied to gold-bearing pyrite products will produce gold-silver doré.

 

The LCT results from the various test programs were used to project flotation metallurgical performance. All of the metal recoveries to the concentrates from the LCTs were adjusted to the target copper grade of the final concentrate. The gold and silver recoveries to the final doré were estimated based on the cyanidation test results including metal losses, which were expected to occur during downstream activated carbon acid washing, desorption and in smelting operations. It should be noted that it is expected the SART circuit should recover approximately 1 to 2% or higher of the total copper from the cyanide leach solution. The additional copper recovery is not included in the copper flotation recovery projection discussed below.

 

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13.6.1Mitchell, East Mitchell, Sulphurets, Upper Kerr, and Upper Iron Cap

 

The results show that the Mitchell mineralization produced better metallurgical performance, compared to the East Mitchell, Sulphurets, Kerr, and Iron Cap mineralization. The metallurgical performance projections of the different types of KSM mineralization are summarized in Table 13.41 to Table 13.44. The estimates are based on a primary grind size of 80% passing approximately 125 µm to 150 µm and a regrind size of 80% passing approximately 20 µm.

 

Table 13.41Cu-Au Flotation Concentrate Grade Versus Cu Head Grade

 

Cu Head Grade
(%)
Cu Concentrate Grade
(%)
Cu Head Grade
(%)
Cu Concentrate Grade
(%)
>0.80 28 0.10-0.15 23
0.40-0.80 26 0.05-0.10 17
0.15-0.40 25 <0.05 5

 

Table 13.42Cu-Au Flotation Concentrate – Metal Recovery Projections

 

Deposit Description Head Grade Recovery
Mitchell Copper Recovery >1.0% Cu = 95%
0.8 – 1.0% Cu = 92%
0.234 – 0.8% Cu = 90.86 x (Cu Head, %) 0.027
0.05 – 0.234% Cu = 18.02 x ln(Cu Head, %) + 113.5
0.02 – 0.05% Cu = 20%
<0.02% = 3%
Gold Recovery n/a = 0.0967 x (Cu Recovery, %) 1.4465 *
Silver Recovery n/a = 1.427 x (Cu Recovery, %) – 70.11

table continues…

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Deposit Description Head Grade Recovery
Mitchell + East Mitchell Upper Copper Recovery >1.0% Cu = 95%
>0.8 – 1.0% Cu = 92%
>0.234 – 0.8% Cu = 90.86 x (Cu Head, %) 0.027
>0.05 – 0.234% Cu = 18.02 x ln(Cu Head, %) + 113.5
0.02 – 0.05% Cu = 20%
<0.02% = 3%
Gold Recovery n/a = 0.0967 x (Cu Recovery, %) 1.4465
Silver Recovery n/a = 1.427 x (Cu Recovery, %) – 70.11
East Mitchell Main Copper Recovery >1.0% Cu = 95%
>0.8 – 1.0% Cu = 92%
>0.205 – 0.8% Cu = 91%
>0.05 – 0.234% Cu = 15.793 x ln(Cu Head, %) + 115.77
0.02 – 0.05% Cu = 20%
<0.02% = 3%
Gold Recovery n/a = 77.324 x ln(Cu Recovery, %) - 287.33
Silver Recovery n/a = 0.1334 x (Cu Recovery, %) + 31.789
Sulphurets Copper Recovery >1.0% Cu = 93%
0.8 – 1.0% Cu = 90%
0.234 – 0.8% Cu = 90.86 x (Cu Head, %) 0.027 – 3.5
0.05 – 0.234% Cu = 18.02 x ln(Cu Head, %) + 110
0.02 – 0.05% Cu = 20%
<0.02% = 3%
Gold Recovery n/a = 52.07 x ln(Cu Recovery, %) – 174.1
Silver Recovery n/a = 1.065 x (Cu Recovery, %) – 44.80; if copper recovery < 50%, use 5%
Upper Kerr Copper Recovery >1.0% Cu = 88%
0.8 – 1.0% Cu = 85%
0.234 – 0.8% Cu = 90.86 x (Cu Head, %) 0.027 – 7
0.05 – 0.234% Cu = 18.02 x ln(Cu Head, %) + 106.5
0.02 – 0.05% Cu = 20%
<0.02% Cu = 3%
Gold Recovery n/a = 171.8 x ln(Cu Recovery, %) – 718; if copper recovery < 70%, use 5%
Silver Recovery n/a = 132.48 x ln(Cu Recovery, %) - 542.9; if copper recovery < 70%, use 5%
table continues…

 

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Deposit Description Head Grade Recovery
Upper Iron Cap Copper Recovery >1.0% Cu = 95%
0.8 - 1.0% Cu = 92%
0.49 - 0.8% Cu = 90%
0.10 - 0.49% Cu = 90.786 x (Cu Head, %) 0.089
0.05 - 0.10% Cu = 30%
<0.05% Cu = 3%
Gold Recovery >2.0 g/t Au = 80%
0.75 – 2.0 g/t Au = 72.5%
0.05 – 0.75 g/t Au = 78.128 x (Au Head, g/t) 0.3012
<0.05 g/t Au = 20
Silver Recovery >20 g/t Ag = 83%
10 – 20 g/t Ag = 78%
0.5 – 10 g/t Ag = 39.945 x (Ag Head, g/t) 0.2602
<0.5 g/t Ag = 5%

Note: ** add 3% additional gold recovery to the copper flotation concentrate for the Mitchell initial pit materials

 

Table 13.43Au-Ag Doré – Cyanide Leach Metal Recovery Projections*

 

Deposit Head Grade Recovery
Mitchell Gold
>10 g/t Au = (98 – ( 0.096 x (Cu Recovery, %) 1.446)) x 80% x 98%
5 – 10 g/t Au = (95 – ( 0.096 x (Cu Recovery, %) 1.446)) x 75% x 98%
0.1 – 5 g/t Au = (87.491 x (Au Head, g/t)0.051 – ( 0.096 x (Cu Recovery, %) 1.446 **)) x 66% x 98%
<0.1 g/t Au = 0%
Silver
>15 g/t Ag = 88 – (1.427 x (Cu Recovery, %) – 70.11)
8- 15 g/t Ag = 86 – (1.427 x (Cu Recovery, %) - 70.11)
1 - 8 g/t Ag = (42.74 x (Ag Head, g/t) 0.336 ) - ( 1.427 x (Cu Recovery, %) - 70.11) ; if <0, use 0%
<1 g/t Ag = 0%
Mitchell + East Mitchell Upper Gold
>10 g/t Au = (98 – (0.096 x (Cu Recovery, %) 1.446)) x 83% x 98%
5 – 10 g/t Au = (95 – (0.096 x (Cu Recovery, %) 1.446)) x 78% x 98%
0.1 – 5 g/t Au = (87.491 x (Au Head, g/t)0.051 – (0.096 x (Cu Recovery, %) 1.446)) x 72% x 98%
<0.1 g/t Au 0
Silver
>15 g/t Ag = 88 – (1.427 x (Cu Recovery, %) – 70.11)
8- 15 g/t Ag = 86 – (1.427 x (Cu Recovery, %) - 70.11)
1 - 8 g/t Ag = (42.74 x (Ag Head, g/t) 0.336 ) - ( 1.427 x (Cu Recovery, %) - 70.11) ; if <0, use 0%
<1 g/t Ag 0
  table continues…

 

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Deposit Head Grade Recovery
East Mitchell Main Gold
>10 g/t Au = (98 – ( 0.096 x (Cu Recovery, %) 1.446)) x 80% x 98%
5 – 10 g/t Au = (95 – ( 0.096 x (Cu Recovery, %) 1.446)) x 70% x 98%
0.1 – 5 g/t Au = (3.5194 x ln(Au Head, g/t) + 87.605 – (77.324 x ln(Cu Recovery, %) - 287.33)) x 55% x 98%
<0.1 g/t Au 0
Silver
>15 g/t Ag = 88 – (1.427 x (Cu Recovery, %) – 70.11)
8- 15 g/t Ag = 86 – (1.427 x (Cu Recovery, %) - 70.11)
1 - 8 g/t Ag = (4.1664 x ln(Ag Head, g/t) + 64.021) - (0.1334 x (Cu Recovery, %) + 31.789)); if <0, use 0%
<1 g/t Ag 0
Sulphurets Gold
>10 g/t Au = (98 – (52.07 x ln(Cu Recovery, %) – 174.1)) x 70% x 98%
5 - 10 g/t Au = (95 – ( 52.07 x ln(Cu Recovery, %) – 174.1)) x 60% x 98%
0.1 - 5 g/t Au = ((87.491 x (Au Head, g/t)0.051 +3)- (52.07 x ln(Cu Recovery, %) - 174.1)) x 49% x 98%
<0.1 g/t Au = 0%
Silver
>15 g/t Ag = 52.7%
8- 15 g/t Ag = 50.7%
1 – 8 g/t Ag = (42.74 x (Ag Head, g/t) 0.336) - (1.065 x (Cu Recovery, %) - 44.80); if <0, use 0%
<1 g/t Ag = 0%
Upper Kerr Gold
>10 g/t Au = (98 – (171.8 x ln(Cu Recovery, %) – 718)) x 75% x 98%
5 - 10 g/t Au = (95 – (171.8 x ln(Cu Recovery, %) – 718)) x 65% x 98%
0.1 - 5 g/t Au = ((87.491 x (Au Head, g/t)0.051 + 8)- (171.8 x ln(Cu Recovery, %) - 718))) x 57% x 98%
<0.1 g/t Au = 0%
Silver
>15 g/t Ag = (88 – (132.48 x ln(Cu Recovery, %) – 542.9))/100; Cap at 88%
8- 15 g/t Ag = (86 – (132.48 x ln(Cu Recovery, %) – 542.9))/100; Cap at 86%
1 - 8 g/t Ag = (21.59 x ln(Ag Head, g/t) + 40.14) - (132.48 x ln(Cu Recovery, %) - 542.9) ; if <0, use 0%
<1 g/t Ag = 0%
table continues…

 

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Deposit Head Grade Recovery
Upper Iron Cap Gold
>2 g/t Au = 8%
0.05 - 2 g/t Au = 10%
<0.05 g/t Au = 5%
Silver
>20 g/t Ag = 8%
10 – 20 g/t Ag = 11%
0.5 – 10 g/t Ag = 16%
<0.5 g/t Ag = 5%

Note: * doré recoveries are in addition to flotation gold and silver recoveries.

** add 3% additional gold recovery to the copper flotation concentrate for the Mitchell initial pit materials

 

Table 13.44Mo Flotation Concentrate Metal Recovery and Grade (Molybdenum Concentrate Grade = 50%)

 

Mo Head Grade
(%)
Mo Recovery
(%)
>0.010 47
0.005-0.010 35
0.0025-0.005 25
<0.0025 0

 

13.6.2Metallurgical Performance Projection – Lower Kerr and Lower Iron Cap

 

After 2016, further metallurgical test work has been conducted, mainly on the mineral samples generated from the lower Kerr and Iron Cap zones. According to the data obtained from the test programs, the copper, gold, and silver recovery performance projections for the lower zones of the Kerr and Iron Cap deposits have been updated and are summarized Table 13.45 to Table 13.47.

 

The similar copper grade for the copper-gold concentrate is assumed for the lower zone mineralization, compared to the other mineralization shown in Table 13.45.

 

The average primary grind size used for the metallurgical performances for both the lower Iron Cap and the lower Kerr lower mineralization is 80% passing approximately 125 µm. For the lower Iron Cap zone mineralization, a diluted slurry solid density of approximately 25% is also considered in the metallurgical performance projections.

 

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Table 13.45Cu-Au Flotation Concentrate –Metal Recovery Projections

 

Deposit Description Head Grade Recovery
Lower Kerr Copper Recovery > 1.5% Cu = 96.5
0.455 – 1.5% Cu = 5.5138 x ln (Cu Head, %) + 93.548
0.20 – 0.455% Cu = 95.434 x (Cu Head, %) 0.0856
0.10 – 0.20% Cu = 78
< 0.10% Cu = 30
Gold Recovery n/a = 0.9812 x (Copper Recovery, %) – 26.327
Silver Recovery n/a = 0.9773 x (Copper Recovery, %) – 33.149; If silver recovery < 0, use 0
Lower Iron Cap Copper Recovery > 1.0% Cu = 95
0.5 – 1.0% Cu = 93
0.2 – 0.5% Cu = 97.915 x (Cu Head, %) 0.0849
0.1 - 0.2% Cu = 106.39 x (Cu Head, %) 0.1371
0.05 - 0.1% Cu = 30
< 0.05% Cu = 3
Gold Recovery > 2.0 g/t Au = 73
0.8 – 2.0 g/t Au = 70
0.15 – 0.8 g/t Au = 64.543 x (Au Head, g/t) 0.021
< 0.15 g/t Au = 40
Silver Recovery >20 g/t Ag = 80
11 – 20 g/t Ag = 78
0.5 – 11 g/t Ag = 46.094 x (Ag Head, %)0.206
<0.5 g/t Ag = 20

 

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Table 13.46Au-Ag Doré – Cyanide Leach Metal Recovery Projections

 

Deposit Head Grade Recovery
Lower Kerr Gold
> 1.0 = 12
0.1 – 1.0 = 14
< 0.1 = 0
Silver
> 4.5 = 11
1 – 4.5 = 16
< 1.0 = 5
Lower Iron Cap Gold
>0.8 g/t Au = 10%
0.15 – 0.8 g/t Au = 11%
<0.15 g/t Au = 7%
Silver
>11 g/t Ag = 8%
0.5 – 11 g/t Ag = 14%
<0.5 g/t Ag = 5%
Notes:1) The doré recoveries are in addition to flotation gold and silver recoveries.

2) Gold and silver recoveries to doré from the flotation products of the first cleaner flotation tailings and the pyrite concentrate are projected based on test work in which the cyanide leaching conditions were not optimized. Whether it is economic to run the cyanide leach circuit on the gold-bearing flotation tailings to recover the additional gold and silver depends on leach circuit feed grades, metal prices, and operating costs. Further test work is recommended to improve metallurgical performances and reduce reagent consumption.

 

The molybdenum metallurgical performance projection is same as shown in Table 13.46 for the lower Iron Cap mineralization; the molybdenum metallurgical performance for the lower Kerr mineralization is summarized in Table 13.47.

 

Table 13.47Mo Flotation Concentrate Metal Recovery and Grade (Molybdenum Concentrate Grade = 50%)

 

Mo Head
(%)
Mo Recovery
(%)
> 0.010; Mo/Cu > 0.006 = 45
0.0050-0.010; Mo/Cu > 0.006 = 35
0.0025-0.0050; Mo/Cu > 0.006 = 25
<0.0025 = 0

 

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14.0Mineral Resource Estimates

 

Mineral Resources for Mitchell and East Mitchell were updated in November of 2021 under the supervision of Henry Kim, P.Geo., of Wood. Mineral Resources for Kerr, Sulphurets, and Iron Cap remain the same as those prepared for the 2020 PFS and were reviewed and endorsed by Henry Kim.

 

A variety of basic descriptive statistics and spatial analyses were completed for each area upon the completion of annual drilling campaigns. These investigations include the generation of grade distribution tables, grade histograms, cumulative probability plots, grade box plots, grade contact plots, down-hole variograms, and directional variograms. In addition, new drill hole results were typically compared against the previous grade model to assess model performance.

 

The following sections summarize the key assumptions, parameters, and methods that were used to estimate Mineral Resources for each of the four deposits for which Mineral Resources have been established.

 

14.1Kerr Deposit

 

The Kerr resource estimate was last updated in 2016 using all drilling data, with the most recent drilling completed in that year. The block model was informed by 223 diamond core holes totaling about 84,770 m of assayed drilling data. Approximately 75% of that drill data was collected by Seabridge. Placer Dome drilled over 80 relatively shallow holes in 1992 while exploring for an open pit copper-gold deposit.

 

14.1.1Grade Distribution – Kerr Deposit

 

The distribution of gold and copper grades within the Kerr deposit were compared with logged lithology and alteration to determine if those attributes could be used for grade estimation purposes. Grade was often seen to cross-cut various lithologic and/or alteration boundaries. For that reason, grade wireframes were designed to constrain the estimate of block grades. Five gold grade wireframes and six copper wireframes were designed by Seabridge’s geological staff.

 

The distribution of gold based on uncapped, uncomposited assay data was summarized at four different cutoff grades by five gold grade wireframes.

 

14.1.2Assay Grade Capping – Kerr Deposit

 

High-grade outlier values were identified using cumulative probability plots for gold, copper, silver, and molybdenum assays. The analysis was completed on the raw assay data prior to compositing for each metal by grade domains.

 

Based on the interpretation of the cumulative probability plots, assays were “cut” or capped at specified thresholds for each metal by grade wireframe domain. Fifty-three gold assays were capped using 1.75 g/t for the lower grade wireframes and 10.0 g/t for the higher grade wireframes. Twenty-three copper assays were capped at 2% for the lower grade domains and 4% for higher grade domains.

 

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14.1.3Drill Hole Compositing – Kerr Deposit

 

The capped drill hole assays were composited into 15 m long composites starting from the drill hole collar keying off of the grade wireframes to start/stop the compositing at wireframe contacts. Two sets of composites were generated, one for precious metals and another for base metals. The precious metal composites keyed off of the gold grade wireframes and the base metal composites keyed off of the copper grade wireframes. Most of the original assay data were in the range of 1.5 to 3.0 m long, with the majority being 2 m long. Based on the scale of the deposit, 15 m long composites were deemed to be an appropriate length for estimating Mineral Resources for Kerr.

 

The assays were composited using MineSight® software. Various geological data were assigned to the 15 m long composites using the majority rule method.

 

14.1.4Geological Constraints - Kerr Deposit

 

Lithological, alteration, structural domains, and metal grade envelopes were constructed for the Kerr deposit by Seabridge and reviewed and accepted by the QP responsible for this section of this Report. Initially these various wireframes were interpreted onto cross sections, which were then reconciled in bench plan prior to building the final wireframe. Leapfrog software was used for generating wireframes for various attributes. The Leapfrog wireframes were modified to account for structural and lithologic criteria.

 

14.1.5Variography – Kerr Deposit

 

Spatial continuity was investigated by generating both grade and indicator correlograms using Sage2001 software. Down-hole correlograms were used to establish nugget effects using 2 m long drill hole composites. Directional correlograms were generated using 15 m long drill hole composites in 30-degree vectors in plan and section. The autofit function in Sage2001 was used to model the 37 directional correlograms.

 

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14.1.6Grade Estimation Parameters – Kerr Deposit

 

The Kerr deposit was modeled using MineSight®, a widely recognized commercial mine engineering software package. The non-rotated model uses NAD83 UTM coordinates with 15 m x 15 m x 15 m blocks. Block gold, silver, copper, and molybdenum grades were estimated by two methods: inverse distance weighting (IDW), and nearest neighbour (NN). Trend planes were used to select eligible drill hole composites for the inverse distance grade models in lieu of traditional search ellipses. Four structural domain wireframes were developed to reflect different mineralized strike/dip regimes which allowed for specific trends to be applied tor each area. Two-pass strategy was used for estimating block gold, silver, copper, and molybdenum grades. A minimum of three composites, a maximum of six composites with no more than two composites per drill hole were used in the grade estimate process. The first estimation pass used a very narrow search in the direction perpendicular to the plunge of mineralization (400 m x 400 m x 20 m). All blocks estimated by the first pass were flagged and not overwritten by the second estimation pass which used a 400 m x 400 m x 80 m search relative to the specified trend plane. Grade envelopes were used as the primary constraint for selecting eligible drill hole composites. Gold wireframes were used in the estimate of gold and silver. Copper wireframes were used in the estimate of copper and molybdenum. The gold grade wireframes were not used as hard boundaries in the estimation process. Drill hole composites from an adjacent lower grade domain were potentially allowed to inform model blocks, provided the composites were located within the trend plane search parameters. For example, a model block within the 0.20 g/t domain could use composites from either the 0.20 g/t domain or the adjacent 0.10 g/t. The highest-grade copper domain (+ 1%) was treated as a hard contact in the estimation process based on visual observations of drill core.

 

14.1.7Grade Model Verification - Kerr Deposit

 

Estimated block grades were verified by visual and statistical methods. Block grades (gold, silver, copper, and molybdenum) were visually compared with drill hole composite grades in cross section and level plan views. Figure 14.1and Figure 14.2 are east-west cross sections through the Kerr block model showing gold and copper grades, respectively. For reference, the location of the Kerr cross sections is illustrated in Figure 10.2, a drill hole plan map for the Kerr deposit. Figure 14.3 and Figure 14.4 are block model level maps drawn at the 850 m elevation through the Kerr block model showing estimated block/composite gold and copper grades, respectively.

 

NN models were prepared for gold, copper, silver, and molybdenum in order to check for potential global biases in the estimated block grades. Swath plots were generated by block column (easting), block row (northing), and block level (elevation) comparing the NN grade with the IDW grade. Figure 14.5 shows two grade swath plots for gold (upper) and copper (lower) that compare the IDW grades in red versus the NN grades in blue by elevation. The results show that the inverse models compare very well with the NN grades on a local basis.

 

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Figure 14.1Kerr Cross Section 6,259,650 N. – Gold

 

 

 

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Figure 14.2Kerr Cross Section 6,259,650 N. – Copper

 

 

 

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Figure 14.3Kerr 850 m Level Plan – Gold

 

 

 

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Figure 14.4Kerr 850 m Level Plan - Copper

 

 

 

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Figure 14.5Kerr Gold-Copper Swath Plots by Elevation

 

 

 

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14.1.8Resource Classification – Kerr Deposit

 

Mineral Resources were defined by drill hole spacing and mineralized continuity. Both Indicated and Inferred wireframes were constructed in cross section and level plan keying on distribution of drilling, mineralized trends, and post mineral structures.

 

Indicated Mineral Resources were restricted to the upper portion of the deposit where the density of drilling is much higher than the deeper mineralization. The Inferred Mineral Resource shape was generated to envelope the Indicated shape and to follow the two steep west-dipping mineralized limbs of the deposit. In general, the drill hole spacing for Indicated Resources is about 50 m while the spacing for Inferred Resources varies between 125 m and 150 m, with wider spacing with depth.

 

14.2Sulphurets Deposit

 

The Sulphurets geological and grade models were updated in mid 2019 using data through the 2018 drilling campaign. After the Sulphurets model was updated, additional drill holes were completed in late 2019 to the west and east of the currently recognized mineralized system. The 2019 drill holes were compared against the EOY2018 model and were found to confirm the geological interpretations and were consistent with the grade model in the area of the 2019 drilling.

 

14.2.1Grade Distribution – Sulphurets Deposit

 

The distribution of gold and copper grades within the Sulphurets deposit were compared with logged lithology and alteration to determine if those attributes could be used for grade estimation purposes. Grade was often seen to cross-cut various lithological and/or alteration boundaries or cut off by structures that form discrete blocks or panels of mineralization. Four gold and copper grade wireframes were designed by Seabridge’s geological staff by structural block. In addition to the grade wireframes, three post-mineral dikes and a distinct monzonite intrusive unit were combined with the grade domains as an aid in constraining the estimate of block grades.

 

The distribution of gold based on uncapped, uncomposited assay data is summarized at four different cutoff grades by gold grade and lithological wireframes. The average uncapped gold grade is 0.44 g/t for all data. A high coefficient of variation exists for all data and for the 0.125 g/t gold envelope was highly skewed by a single gold assay grade of 1,580 g/t from drill hole S-18-81. This highly anomalous sample was capped at 2.0 g/t which lowered the coefficient of variation from 23.4 to 1.6. The average copper grade for all data is 0.12% with a coefficient of variation of 1.49. The two higher grade domains (0.20% and 0.50%) have coefficients of variation that are less than 1.0, reflecting the role of grade domaining in reducing variability.

 

14.2.2Assay Grade Capping – Sulphurets Deposit

 

High-grade outlier values were identified using cumulative probability plots for gold, copper, silver, and molybdenum assays. The analysis was completed on the raw assay data prior to compositing for each metal by grade domains. Based on the interpretation of the cumulative probability plots, assays were “cut” or capped at specified thresholds for each metal by grade wireframe domain. A total of 112 gold assay samples were capped. The lower grade gold domains, post mineral dikes, and monzonite were capped at 1 g/t. The upper grade gold domains (0.25 to 0.50 and + 0.50 g/t domains) were capped at 5.0 and 10.0 g/t, respectively. 137 copper assays were capped at 0.60% and 1.00% for the lower grade domains and 2% for high-grade domain (e.g. + 0.5%).

 

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14.2.3Drill Hole Compositing – Sulphurets Deposit

 

The capped drill hole assays were composited into 10 m long composites starting from the drill hole collar keying off of the grade wireframes to start/stop the compositing at wireframe contacts using MineSight® software. Various geological data were assigned to the drill hole composites using the majority rule method. Four sets of composites were generated, one for each of the estimated metals (i.e. gold, silver, copper, and molybdenum). Most of the original assay data were in the range of 1.5 to 3.0 m long, with the majority being 2 m long. Based on the scale of the deposit and model block dimensions, 10 m long composites were deemed to be an appropriate length for estimating Mineral Resources.

 

14.2.4Geological Constraints – Sulphurets Deposit

 

Lithological, alteration, structural domains, and metal grade envelopes were constructed for the Sulphurets deposit by Seabridge and reviewed and accepted by the QP responsible for this section of this Report. Leapfrog software was used to generate preliminary wireframes that were further modified to account for various structural and lithological criteria.

 

A review of gold and copper distribution by lithology and alteration determined that independently constructed grade wireframes are more appropriate for estimating block grades. Five structural domain wireframes were also used in the interpolation plan as hard boundaries along post mineral structures that locally juxtapose well mineralized material against poorly mineralized rock.

 

14.2.5Variography – Sulphurets Deposit

 

Spatial continuity was investigated by generating both grade and indicator correlograms using Sage2001 software. Down-hole correlograms were used to establish nugget effects using 2 m long drill hole composites. Directional correlograms were generated using 10 m long drill hole composites in 30-degree vectors in plan and section. The autofit function in Sage2001 was used to model the 37 directional correlograms.

 

14.2.6Grade Estimation Parameters – Sulphurets Deposit

 

The Sulphurets deposit was modeled using MineSight®, a widely recognized commercial mine engineering software package. The non-rotated model uses NAD83 UTM coordinates with 12.5 m x 12.5 m x 15m blocks.

 

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Block gold, copper, silver and molybdenum grades were estimated by three primary methods: IDW, Ordinary Kriging (OK), NN. The OK model was selected for resource declaration after comparing that model with the IDW model.

 

Independently constructed gold, copper, silver, and molybdenum wireframes provided the principal constraint in the grade estimation plan. In addition to the grade envelopes, five structural domain wireframes were developed to reflect different mineralized strike/dip regimes which allowed for specific search ellipses to be incorporated for each structural block. The grade shells were treated as soft contacts in the grade estimation plan where lower grade samples were allowed to inform the next highest grade shell but the higher grade samples were not allowed to inform adjacent lower grade zones.

 

A two-pass strategy was used for estimating block gold, silver, copper, and molybdenum grades for both the IDW and OK models. The first estimation pass used a 275 m x 275 m x 55 m search ellipse that required a minimum of five drill hole composites, a maximum of nine composites with no more than three composites per drill hole. That criteria resulted in blocks being estimated at least two drill holes. All blocks estimated by the first pass were flagged and not overwritten by the second estimation pass which used a 125 m x 125 m x 25 m search that required a minimum of two drill hole composites, a maximum of nine composites with no more than three composites per drill hole. That criteria resulted in blocks that could have been estimated by one or more drill holes.

 

Narrow, post-mineral dike wireframes were developed for the Sulphurets geological model. The volume percentage of these narrow dikes was stored in the block model. Two grades were estimated for blocks that contained less than 50% dike material. One grade was based on dike-only samples, and the other block grade was estimated with samples that matched the grade shell code of the remaining majority percentage of the block. A final weighted block grade was calculated using the two block proportions and their associated grades. Blocks that contained more than 50% dike were considered to be 100% dike material.

 

14.2.7Model Validation – Sulphurets Deposit

 

Estimated block grades (gold, silver, copper, and molybdenum) were compared to drill hole composite grades in cross section and level plan views. Figure 14.6 and Figure 14.7 are northwest-southeast cross sections drawn through the Sulphurets block model (refer to Figure 10.3 for the location of the Cross Section 20). Figure 14.8 and Figure 14.9 are block model level maps drawn at the 1135 m elevation through the Sulphurets block model showing estimated block grades and drill hole composite for gold and copper, respectively.

 

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Figure 14.6Sulphurets Cross Section 20 - Gold

 

 

 

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Figure 14.7Sulphurets Cross Section 20 – Copper

 

 

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Figure 14.8Sulphurets 1135m Level Plan - Gold

 

 

 

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Figure 14.9Sulphurets 1135 m Level Plan – Copper

 

 

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The grade models were also validated by comparing the OK block grades against NN models that were generated for gold, copper, silver, and molybdenum.

 

Grade swath plots were generated for rows (east-west), columns (north-south) and levels (elevations) through the block model comparing the OK and NN models at a zero cutoff grade. Figure 14.10 shows swath plots for gold (upper) and copper (lower) by elevation. These swath plots show Indicated and Inferred Resource grades.

 

Figure 14.10Sulphurets Gold-Copper Swath Plots by Elevation

 

 

 

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14.2.8Resource Classification – Sulphurets Deposit

 

Mineral Resources were assigned to the estimated blocks by constructing 3D solids for Indicated and Inferred Mineral Resources. These shapes were based on mineralized continuity as defined by exploration drill hole results. The average drill hole spacing for Indicated and Inferred blocks is approximately 50 m to 75 m and 75 m to 125 m, respectively.

 

14.3Mitchell and East Mitchell Deposit

 

Mitchell resource model, including East Mitchell, was updated in late 2021. That model was constructed using drilling data collected through 2021.

 

14.3.1Metal Distribution – Mitchell and East Mitchell Deposit

 

Basic descriptive statistics were generated by lithology, structural block, and deposit. Six gold and copper estimation domains were constructed by grouping lithology domains. Table 14.1 summarizes uncapped gold assay distributions by lithology wireframes. Table 14.2 summarizes uncapped copper assay distributions by lithology wireframes.

 

Table 14.1Distribution of Gold by Lithology and Estimation Domain – Mitchell and East Mitchell Deposit

 

Estimation
Domain
Geology

Geology
Code

Number
of Data
Au (g/t) CV
Mean Maximum Minimum
EDOM 2 SEDS 10 24,415 0.287 26.500 0.001 1.279
VOLC 20 14,450 0.672 17.650 0.003 1.152
P2 30 21,084 0.491 31.500 0.003 1.046
EDOM 1 P3 40 980 0.828 12.500 0.067 0.814
H1B 50 2,253 0.666 6.450 0.005 0.604
P1B 60 4,224 0.761 6.830 0.003 0.574
H1 70 2,611 0.763 6.420 0.066 0.528
P1 80 7,008 0.804 53.800 0.001 1.083
EDOM 5 QABX 90 635 0.300 2.620 0.001 0.916
EDOM 4 P5 100 260 0.067 4.980 0.015 4.737
P8 110 1,851 0.062 3.030 0.003 2.424
EDOM 3 K2 120 1,849 0.499 4.700 0.011 0.725
K3 130 39 0.480 2.730 0.050 0.873
K4 140 412 0.332 2.900 0.030 1.123

Source: (Wood, 2021)

Note: CV = coefficient of variation

 

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Table 14.2Distribution of Copper by Lithology and Estimation Domain – Mitchell and East Mitchell Deposit

 

Estimation
Domain
Geology Geology
Code
Number
of Data
Cu (%) CV
Mean Maximum Minimum
EDOM 2 SEDS 10 24,415 0.069 3.510 0.000 1.135
VOLC 20 14,450 0.057 4.500 0.000 3.126
P2 30 21,084 0.133 17.800 0.000 1.214
EDOM 1 P3 40 980 0.199 0.933 0.005 0.484
H1B 50 2,253 0.146 0.512 0.001 0.351
P1B 60 4,224 0.201 1.060 0.001 0.556
H1 70 2,611 0.157 0.631 0.005 0.316
P1 80 7,008 0.184 4.090 0.002 0.499
EDOM 5 QABX 90 635 0.300 3.640 0.003 1.402
EDOM 4 P5 100 260 0.102 0.574 0.013 0.657
P8 110 1,851 0.061 0.852 0.001 1.144
EDOM 3 K2 120 1,849 0.171 0.913 0.001 0.641
K3 130 39 0.010 0.039 0.001 0.926
K4 140 412 0.084 0.836 0.000 1.224

Source: (Wood, 2021)

 

14.3.2Assay Grade Capping – Mitchell Deposit

 

Cumulative probability plots were used to identify high-grade outliers for gold, copper, silver, and molybdenum based on the original assay samples. No capping was applied for gold and molybdenum. However, seven strongly anomalous silver assays were capped at 300 g/t Ag and one anomalous copper assay grade was capped at 5% prior to calculation of downhole composites. Table 14.3 presents capped composites and their capped values.

 

Table 14.3Capped Silver Assay Intervals and Grades

 

Hole From To Original Assay
Grade
Capped Grade
M-10-115 141 142 1,630 g/t Ag 300 g/t Ag
M-10-115 141 142 17.8% Cu 5.0 % Cu
MZ-091 93 94.5 640 g/t Ag 300 g/t Ag
M-11-127 949 951 591 g/t Ag 300 g/t Ag
M-18-140 430 432 486 g/t Ag 300 g/t Ag
MZ-091 412 413 442 g/t Ag 300 g/t Ag
M-21-151 321.05 322.5 408 g/t Ag 300 g/t Ag
M-11-126 360 362 354 g/t Ag 300 g/t Ag

Source: (Wood, 2021)

 

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14.3.3Drill Hole Compositing – Mitchell Deposit

 

Drill hole assay data (both uncapped and capped intervals) were composited into 7.5 m long composites starting from the drill hole collar. Most of the original assay data were in the range of 1.5 m to 3 m long, with the majority being 2 m long. Based on the scale of the deposit, 7.5 m long composites were deemed to be an appropriate length for estimating Mineral Resources.

 

After compositing, block model attributes like lithology, structural block, and deposits were backtagged from the model to the drill hole composites.

 

14.3.4Variography – Mitchell and East Mitchell Deposit

 

A variety of grade and indicator variograms were generated for the Mitchell deposit using MineSight® and Sage2001 software. Down-hole correlograms for Mitchell and Mitchell East based on 7.5 m long drill hole composites.

 

Modelled gold (upper) and copper (lower) correlograms 7.5 m drill hole composites for Mitchell and East Mitchell. Nested spherical models were used in modeling the Mitchell and East Mitchell gold and copper correlograms. The gold and particularly the copper directional correlograms show relatively long ranges.

 

14.3.5Grade Estimation Parameters – Mitchell Deposit

 

Gold, copper, silver, and molybdenum block grades were estimated for the Mitchell deposit using a multi-pass OK method. Grades were estimated using 7.5 m long drill hole composites. OK was selected as the approach for gold, copper, silver, and molybdenum estimation due to the moderate log-normal distribution of gold, copper, silver, and molybdenum grades. Light outlier restriction is used to control against over-projection of outlier grades in areas of relatively sparse drilling. The outlier restriction grade thresholds were selected individually for each estimation domain and metal, and the range was set to 80 m so that over-threshold composites were not used to estimate blocks beyond a radius of 80 m. Estimation runs were set up for each metal, site, and estimation domain to allow for separate configuration of search ellipsoid orientation and range, variogram model, outlier restriction parameters. Composite selection parameters were refined iteratively to produce a grade tonnage curve and change of support approaching a theoretical HERCO grade tonnage curve. Table 14.4 lists the estimation parameters used for gold, copper, silver, and molybdenum grade estimation.

 

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Table 14.4Estimation Parameters for Gold, Copper, Silver and Molybdenum

 

EDOM 1 2 3 4 5 1 2 3 4 5 1 2 3 4 5 1 2 3 4 5
Run/Rpt file name G11 G12 G13 G14 G15 C11 C12 C13 C14 C15 S11 S12 S13 S14 S15 M11 M12 M13 M14 M15
Block Variable AUOK AUOK AUOK AUOK AUOK CUOK CUOK CUOK CUOK CUOK AGOK AGOK AGOK AGOK AGOK MOOK MOOK MOOK MOOK MOOK
Comp Variable AU AU AU AU AU CU CU CU CU CU AG AG AG AG AG MO MO MO MO MO
Max search dist. 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400
Min composites 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3
Max composites 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8
Max comp/hole 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2
Major-axis (Y) dist. 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400
Semi-major (X) dist. 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300
Minor (Z) distance 100 100 100 100 100 100 100 100 100 100 100 100 100 100 100 100 100 100 100 100
Rotation in Z (LH) 345 345 345 345 345 345 345 345 345 345 345 345 345 345 345 345 345 345 345 345
Rotation in X’ (RH) -45 -45 -45 -45 -45 -45 -45 -45 -45 -45 -45 -45 -45 -45 -45 -45 -45 -45 -45 -45
Rotation in Y’ (LH) 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0
Outlier Threshold 1.5 1 1 1 1 1 0.4 0.4 0.4 1.5 12 30 30 4 30 0.03 0.02 0.03 0.02 0.04
OLR Distance 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80
C0 0.2 0.2 0.2 0.2 0.2 0.2 0.2 0.2 0.2 0.2 0.4 0.4 0.4 0.4 0.4 0.2 0.2 0.2 0.2 0.2
C1 0.35 0.35 0.35 0.35 0.35 0.3 0.3 0.3 0.3 0.3 0.2 0.2 0.2 0.2 0.2 0.3 0.3 0.3 0.3 0.3
A1 Major 140 140 140 140 140 120 120 120 120 120 130 130 130 130 130 80 80 80 80 80
A1 Semi-major 140 140 140 140 140 120 120 120 120 120 130 130 130 130 130 120 120 120 120 120
A1 Minor 140 140 140 140 140 120 120 120 120 120 130 130 130 130 130 160 160 160 160 160
C2 0.45 0.45 0.45 0.45 0.45 0.5 0.5 0.5 0.5 0.5 0.4 0.4 0.4 0.4 0.4 0.5 0.5 0.5 0.5 0.5
A2 Major 800 800 800 800 800 1200 1200 1200 1200 1200 1000 1000 1000 1000 1000 900 900 900 900 900
A2 Semi-major 1000 1000 1000 1000 1000 560 560 560 560 560 505 505 505 505 505 550 550 550 550 550
A2 Minor 460 460 460 460 460 390 390 390 390 390 745 745 745 745 745 330 330 330 330 330
table continues…

 

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Site 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Un-OLR Run/Rpt G31 G32 G33 G34 G35 C31 C32 C33 C34 C35 S31 S32 S33 S34 S35 M31 M32 M33 M34 M35
Uncapped variable AUUC AUUC AUUC AUUC AUUC CUUC CUUC CUUC CUUC CUUC AGUC AGUC AGUC AGUC AGUC MOUC MOUC MOUC MOUC MOUC

 

EDOM 1 2 3 4 6 1 2 3 4 6 1 2 3 4 6 1 2 3 4 6
Run/Rpt file name G21 G22 G23 G24 G26 C21 C22 C23 C24 C26 S21 S22 S23 S24 S26 M21 M22 M23 M24 M26
Block Variable AUOK AUOK AUOK AUOK AUOK CUOK CUOK CUOK CUOK CUOK AGOK AGOK AGOK AGOK AGOK MOOK MOOK MOOK MOOK MOOK
Comp Variable AU AU AU AU AU CU CU CU CU CU AG AG AG AG AG MO MO MO MO MO
Max search dist. 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400 400
Min composites 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3
Max composites 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8
Max comp/hole 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2
Major-axis (Y) dist. 400 400 400 400 300 400 400 400 400 300 400 400 400 400 300 400 400 400 400 300
Semi-major (X) dist. 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300 300
Minor (Z) distance 200 200 200 200 200 200 200 200 200 200 200 200 200 200 200 200 200 200 200 200
Rotation in Z (LH) 330 330 330 330 60 330 330 330 330 60 330 330 330 330 60 330 330 330 330 60
Rotation in X’ (RH) -50 -50 -50 -50 0 -50 -50 -50 -50 0 -50 -50 -50 -50 0 -50 -50 -50 -50 0
Rotation in Y’ (LH) 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0
Outlier Threshold 1.5 1 1 1 4 1 0.4 0.4 0.2 0.2 12 30 30 4 4 0.03 0.02 0.03 0.02 0.04
OLR Distance 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80 80
C0 0.2 0.2 0.2 0.2 0.2 0.2 0.2 0.2 0.2 0.3 0.35 0.35 0.35 0.35 0.3 0.2 0.2 0.2 0.2 0.4
C1 0.35 0.35 0.35 0.35 0.2 0.4 0.4 0.4 0.4 0.3 0.45 0.45 0.45 0.45 0.2 0.65 0.65 0.65 0.65 0.15
A1 Major 80 80 80 80 50 450 450 450 450 100 140 140 140 140 65 290 290 290 290 80
A1 Semi-major 155 155 155 155 50 280 280 280 280 100 120 120 120 120 80 125 125 125 125 190

table continues…

 

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A1 Minor 285 285 285 285 50 570 570 570 570 400 180 180 180 180 45 280 280 280 280 50
C2 0.45 0.45 0.45 0.45 0.6 0.4 0.4 0.4 0.4 0.4 0.2 0.2 0.2 0.2 0.5 0.15 0.15 0.15 0.15 0.45
A2 Major 1000 1000 1000 1000 430 900 900 900 900 700 1000 1000 1000 1000 380 790 790 790 790 310
A2 Semi-major 450 450 450 450 270 450 450 450 450 1000 330 330 330 330 280 785 785 785 785 190
A2 Minor 820 820 820 820 250 600 600 600 600 450 750 750 750 750 280 765 765 765 765 270
Site 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2
Un-OLR Run/Rpt G41 G42 G43 G44 G46 C41 C42 C43 C44 C46 S41 S42 S43 S44 S46 M41 M42 M43 M44 M46
Uncapped variable AUUC AUUC AUUC AUUC AUUC CUUC CUUC CUUC CUUC CUUC AGUC AGUC AGUC AGUC AGUC MOUC MOUC MOUC MOUC MOUC

Source: (Wood, 2022)

 

NN grade models were generated simultaneously with the OK models using the same constraint requirements. Results from the NN models were used in verifying that the grade model was globally unbiased.

 

14.3.6Grade Model Validation – Mitchell Deposit

 

Estimated block grades (gold, silver, copper, and molybdenum) were compared to drill hole composite grades in cross section and level plan views. Figure 14.11 to Figure 14.14 are block model level maps drawn at the 750 m (Mitchell), 950 m, and 1,550 m (East Mitchell) elevations showing estimated block grades and drill hole composite grades for gold and copper, respectively (refer to Figure 10.4 for the locations of cross section lines).

 

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Figure 14.11Mitchell Cross Section of Gold (upper) and Copper (lower) Assay Composite and Block Grade Estimates

 

Source: (Wood, 2022)

 

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Figure 14.12East Mitchell Cross Section of Gold (upper) and Copper (lower) Assay Composite and Block Grade Estimates

 

 

 

Source: (Wood, 2022)

 

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Figure 14.13Mitchell 750 m El Bench Plan of Gold (upper) and Copper (lower) Assay Composite and Block Grade Estimates

 

 

 

Source: (Wood, 2022)

 

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Figure 14.14950 m El and 1550 m El Bench Plan of Gold (upper) and Copper (lower) Assay Composite and Block Grade Estimates

 

 

 

Source: (Wood, 2022)

 

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The NN validation model was built using two 7.5 m assay composites to match the 15 m block height of the grade estimates and selective mining unit (SMU). The NN model statistics reflect the de-clustered assay statistics and are a good basis to check for global bias in the kriged grade estimates for gold, copper, silver, and molybdenum. Global bias was calculated by expressing the difference of the unrestricted kriged grade versus the nearest neighbor validation model grade. A target of less than 2% difference was set and achieved for gold, copper, and molybdenum grade estimates for Measured and Indicated (M&I) blocks in EDOM 1 and EDOM 6, the high-grade domains. The biases of the Measured, Indicated, and Inferred (MII) blocks are also within tolerance for gold, copper, silver, and molybdenum grades for these domains. Biases are slightly greater for the other lower grade domains but the unrestricted grade estimates are within 5% of the nearest neighbor validation model for all grades and estimation domains, with the exception of EDOM 3 (P5, P8) and EDOM 5 (QABX) for silver and EDOM 5 for molybdenum which have little influence on block value or overall tonnage in the Mitchell-East Mitchell Mineral Resource estimate.

 

Metal reduction was also calculated by expressing the difference of the outlier restricted and unrestricted block estimates as a percentage. Total metal reduction for gold and copper in M&I blocks is 1.5% and 0.4% respectively. Metal reduction for gold in MII blocks increase to 2.1 % as the distance from Inferred blocks to drillhole composites is greater than for M&I blocks and the influence of the 80 m restriction distance increases. Metal reduction for silver is greater, totaling 2.8% for M&I blocks and 3% for MII blocks as more aggressive grade thresholds were set for silver compared to the other grades in order to control the influence of the more important higher grade tail in the silver grade distribution.

 

Table 14.5Validation of Global Bias and Metal Reduction

 

Gold EDOM Tonnage (Mt) AuOK (g/t) Metal
Red. (%)
Global
Bias vs
NN (%)
Tonnage (Mt) AuOK (g/t) Metal
Red. (%)
Global
Bias vs
NN (%)
P1, P1B/H1, H1B/P3 1 937 0.73 -1.4 -0.3 1,002 0.72 -1.7 0.2
SEDS, VOLC, P2 2 4,725 0.35 -1.4 3.0 9,723 0.27 -1.7 3.7
P5, P8 3 210 0.44 -3.9 0.7 263 0.43 -4.1 0.5
K3, K4 4 0 0.00 -- -- 536 0.06 -7.6 -0.4
QABX 5 91 0.31 -7.2 3.5 135 0.30 -7.6 2.7
Gold Zone 6 151 0.90 0.2 0.6 154 0.89 0.2 0.6
Total -- 6,114 0.43 -1.5 2.4 11,812 0.31 -2.1 3.1
table continues…

 

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Copper EDOM Tonnage
(Mt)
CuOK (%) Metal
Red. (%)
Global
Bias vs
NN (%)
Tonnage
(Mt)
CuOK (%) Metal
Red. (%)
Global
Bias vs
NN (%)
P1, P1B/H1, H1B/P3 1 937 0.18 0.1 -0.1 1,002 0.18 0.1 0.5
SEDS, VOLC, P2 2 4,725 0.08 -0.4 0.8 9,724 0.07 -0.3 1.4
P5, P8 3 210 0.16 -1.7 0.5 263 0.16 -2.0 0.3
K3, K4 4 0 0.00 -- -- 536 0.07 0.0 -3.3
QABX 5 91 0.26 -1.4 3.1 135 0.23 -1.3 5.8
Gold Zone 6 151 0.04 0.0 -1.8 154 0.04 0.0 -1.8
Total -- 6,114 0.10 -0.4 0.7 11,813 0.09 -0.3 1.1
Silver EDOM Tonnage
(Mt)
AgOK (g/t) Metal
Red. (%)
Global
Bias vs
NN (%)
Tonnage (Mt) AgOK (g/t) Metal
Red. (%)
Global
Bias vs
NN (%)
P1, P1B/H1, H1B/P3 1 937 2.5 -3.3 -1.9 1,002 2.5 -3.5 -1.5
SEDS, VOLC, P2 2 4,725 1.9 -2.5 0.5 9,724 1.8 -2.7 -0.3
P5, P8 3 210 3.0 -9.3 2.9 263 2.8 -8.2 3.2
K3, K4 4 0 0.0 -- -- 536 0.8 -4.2 -7.1
QABX 5 91 5.1 -5.4 8.7 135 4.6 -5.5 10.7
Gold Zone 6 151 1.3 -0.1 -0.7 154 1.3 -0.1 -0.4
Total   6,114 2.07 -2.8 0.3 11,813 1.87 -3.0 -0.5
Molybdenum EDOM Tonnage
(Mt)

MoOK

 

(%)

 

Metal
Red. (%)
Global
Bias vs
NN (%)
Tonnage
(Mt)

MoOK

 

(%)

 

Metal
Red (%)
Global
Bias vs
NN (%)
P1, P1B/H1, H1B/P3 1 937 0.007 0.0 0.0 1,002 0.006 0.0 0.0
SEDS, VOLC, P2 2 4,719 0.006 -1.7 1.7 9,693 0.005 0.0 2.1
P5, P8 3 210 0.005 -6.3 -5.9 263 0.004 -10.4 -4.0
K3, K4 4 0 0.000 -- -- 536 0.002 0.0 0.0
QABX 5 91 0.008 -10.2 -1.1 135 0.007 -9.8 -3.5
Gold Zone 6 151 0.011 0.0 0.0 154 0.011 0.0 0.0
Total -- 6,108 0.01 -1.7 1.1 11,782 0.00 -0.3 1.6
Source: (Wood, 2022)                          

 

Swath plots were constructed for gold, copper, silver, and molybdenum grades for M&I blocks at Mitchell and East Mitchell (Figure 14.15 and Figure 14.16). High gold and copper grades at the core and trends to lower grades at the edges of the Mitchell deposit are evident in estimated grade trends. The trends of decreasing silver and increasing molybdenum grades with depth are also evident. The high gold grade and lower copper and silver grades of the gold zone at the top of the East Mitchell deposit is evident on the East Mitchell swath plots. The influence of the tight drilling in the higher grade core of the East Mitchell deposit can be seen in the grade profiles and histograms of assay composite data compared to the block estimates.

 

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Figure 14.15Swath Plots of Gold, Copper, Silver, and Molybdenum grades for M&I Blocks at Mitchell

 

Source: (Wood, 2022)

 

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Figure 14.16Swath Plots of Gold, Copper, Silver, and Molybdenum grades for M&I Blocks at East Mitchell

 

 

Source: (Wood, 2022)

 

The gold grade-tonnage curve and change of support from assay composite support to SMU support was validated by adjusting the declustered assay composite distribution to the theoretical SMU using a variance correction factor derived from the block dispersion variance (BDV) for the 25 m x 25 m x 15 m SMU and the unit-sill correlogram models for gold at Mitchell and East Mitchell.

 

Grade models were initially estimated with a maximum of three 7.5 m assay composites per drillhole and a total maximum of 12 assay composites, but initial HERCO validation indicated that grade tonnage curves were too smooth, so runs with maximum two composites per hole, maximum ten composites, and maximum eight composites were executed to reduce the smoothing of the kriged models. Table 14.6 lists the CVs of the final block grade gold estimates (OK CV) using a maximum of two 7.5 m assay composites per drillhole and a maximum of eight assay composites with the nearest neighbor block grades (NN CV), the BDV, and the target CV for the 25 m x 25 m x 15 m SMU.

 

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Table 14.6Change of Support for Gold at Mitchell and East Mitchell (Main Zone)

 

Area SMU OK CV NN CV BDV SMU CV
Mitchell 25x25x15 0.62 0.77 0.73 0.57
East Mitchell 25x25x15 0.75 0.94 0.73 0.68

 

Source: (Wood, 2022)

 

The global variance of the OK gold grade estimates and the Mitchell OK copper estimates have 5% to 10% greater CV, and the East Mitchell OK copper variance is lower than the target grade variance calculated using the BDV from the unit-sill correlogram models for both metals and deposits. However, a HERCO grade-tonnage curve indicates that the grade and tonnage above cut off for gold and copper (Figure 14.17 and Table 14.7) show that the OK grade estimates for M&I blocks are only about 2% smooth with the OK block grades reporting 2.1% higher tonnage at 1.1% lower grade than the HERCO validation curve above a cutoff grade of 0.2 g/t Au. The grade-tonnage curve for M&I blocks at East Mitchell is also about 2% smoother than the theoretical HERCO grade tonnage curve (Figure 14.18 and Table 14.8).

 

Figure 14.17Grade Tonnage Curve for Mitchel Gold Estimates and Theoretical HERCO Curve

 

 

Source: (Wood, 2022)

 

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Table 14.7Grade-Tonnage Metrics for Mitchell Gold Grade Estimates and Theoretical HERCO Distribution

 

Difference (Estimate-HERCO, %)
Cutoff (Au, g/t) Tonnage (%) Grade (%) Metal (%)
0 0.0 0.0 0.0
0.1 -1.5 1.4 -0.1
0.2 2.1 -1.1 1.0
0.3 3.7 -2.0 1.6
0.4 5.2 -3.0 2.1
0.5 5.5 -3.8 1.5
0.6 5.8 -5.0 0.5
0.7 3.0 -6.4 -3.6
0.8 -7.9 -7.1 -14.5
0.9 -27.1 -6.5 -31.8

 

Source: (Wood, 2022)

 

Figure 14.18Grade Tonnage Curve for East Mitchell Gold Estimates and Theoretical HERCO Curve

 

 

Source: (Wood, 2022)

 

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Table 14.8Grade-Tonnage Metrics for East Mitchell Gold Grade Estimates and Theoretical HERCO Distribution

 

Difference (Estimate-HERCO, %)
Cutoff (Au, g/t) Tonnage (%) Grade (%) Metal (%)
0 0.0 0.0 0.0
0.1 3.0 -2.5 0.4
0.2 2.1 -2.2 -0.2
0.3 2.1 -2.4 -0.4
0.4 1.7 -2.8 -1.1
0.5 -1.0 -2.4 -3.4
0.6 -2.1 -2.7 -4.8
0.7 -6.4 -2.3 -8.5
0.8 -11.0 -1.6 -12.4
0.9 -15.7 -0.6 -16.2

 

Source: (Wood, 2022)

 

14.3.7Density Estimation

 

A boxplot showing the statistics of density measurements for each lithology unit in the Mitchell-East Mitchell dataset is shown in Figure 14.19. The ranges of density values is relatively small for the Mitchell and East Mitchell deposits and coefficients of variation of density values by lithological unit are less than 0.05. The SEDS (10), VOLC (20), and altered and mineralized porphyry and hybrid units all have a mean density of 2.77 to 2.82 t/m3. Density was estimated by interpolating density determinations using inverse distance weighting to the second power.

 

Figure 14.19Boxplot of Density Measurements by Lithology

 

 

Source: (Wood, 2022)

 

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14.3.8Resource Classification – Mitchell Deposit

 

The definition of the Indicated confidence category are in accordance with CIM Definition Standards (2014), where estimates with Indicated confidence are derived from adequately detailed and reliable data sufficient to assume continuity of geology and grade between drillhole intersections and estimates have sufficient confidence to support mine planning and deposit evaluation. Guidelines for Measured confidence are those estimates derived from data with sufficiently tight drill spacing to confirm continuity of geological features and grade, and estimates can be used for detailed mine planning. The approach for resource confidence classification for the 2021 Mitchell-East Mitchell estimate is to take into account the scale of the geological features in the 2021 lithological model that are used in the construction of estimation domains, and the continuity of gold grades and estimation error calculated. Copper, silver, and molybdenum grades have a relatively minor contribution to block value and were not included in the assessment of grade continuity for confidence classification.

 

A drillhole spacing study was carried out to calculate the error of estimation of gold grades for a monthly production panel based on a nominal production throughput rate of 100 ktpd and panel measuring 160 m x 160 m x 45 m. The CV of gold grades and EDOM1 gold grade correlogram models for Mitchell and East Mitchell were used to calculate the kriging variance (KV) for the monthly panel which was used to calculate estimation error at the 90th confidence interval (CI) for quarterly and annual production volumes according to the formula:

 

Quarterly Error (90th CI) = 1.645*(KV^0.5)*CV/(3^0.5)

 

Annual Error (90th CI) = 1.645*(KV^0.5)*CV/(12^0.5)

 

The results of the study are similar for Mitchell and East Mitchell, and indicate that precise estimates of gold grades for annual production volumes are estimated with relative confidence due to the relatively low coefficient of variation of gold grades, the relatively long range of continuity of gold grades, and the large volume estimated. However, a drill spacing of 100 m x 100 m is required to achieve estimates of approximately +/-15% at the 90th CI (Table 14.9).

 

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Table 14.9Drillhole Spacing Study Results for Mitchell and East Mitchell

 

Mitchell   Measured Indicated
      Quarterly Error Annual Error
  KV CV 90th CI 90th CI
100 m x 100 m 0.0384 0.933 17.4 -17.4 8.7 -8.7
150 m x 150 m 0.0609 0.933 21.9 -21.9 10.9 -10.9
200 m x 200 m 0.1328 0.933 32.3 -32.3 13.3 -13.3
East Mitchell   Measured Indicated
      Quarterly Error Annual Error
  KV CV 90th CI 90th CI
100 m x 100 m 0.0298 0.933 15.3 -15.3 7.6 -7.6
150 m x 150 m 0.0595 0.933 21.6 -21.6 10.8 -10.8
200 m x 200 m 0.1338 0.933 32.4 -32.4 13.3 -13.3
               

Criteria for classification of M&I Mineral Resources are as follows:

 

A drill pattern of less than 200 m is required to define the narrow late dykes and the shape of the P1/H1 and P1B/H1B units that make up the high-grade estimation domain. A distance of 125 m was selected as the maximum average distance to the nearest two holes to define Indicated Mineral Resources.

 

Additional infill drilling is required to confirm the geological continuity and shape of the mineralized zone and reduce estimation error to 10%-15% at the 90th CI for detailed mine planning. A maximum distance of 80 m between two holes was selected as the limit of drill spacing required to produce resource estimates of Measured confidence.

 

Blocks within 200 m of two drillholes have been identified as having sufficient support to produce estimates of Inferred confidence.

 

14.4Iron Cap Deposit

 

The Iron Cap grade model was last updated in 2018 and is supported by 99 diamond drill holes totaling about 66,740 m and is based on OK methods that used conventional search ellipses.

 

14.4.1Grade Distribution – Iron Cap Deposit

 

Basic descriptive statistics were generated for various logged attributes like lithology and alteration. The interpretation of those results did not identify any single or combined attributes that adequately defined the principal controls of mineralization. Independently generated gold, copper, and molybdenum wireframes were developed by Seabridge’s geological staff.

 

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14.4.2Assay Grade Capping – Iron Cap Deposit

 

Cumulative probability plots were used to identify high-grade outliers for gold, copper, silver, and molybdenum based on the original assay samples. The assays were initially examined with respect to logged lithology and alteration types. However, the final analysis of high-grade outliers was completed using the grade wireframes that were used to interpolate block grades.

 

14.4.3Drill Hole Compositing – Iron Cap Deposit

 

The capped drill hole assays were composited into 10 m long composites starting from the drill hole collar keying off of the grade wireframes to start/stop the compositing at wireframe contacts using MineSight® software. Various geological data were assigned to the drill hole composites using the majority rule method. Two sets of composites were generated, one for precious metals (gold and silver) and one for base metals (copper and molybdenum). Most of the original assay data were in the range of 1.5 to 3.0 m long, with the majority being 2 m long. Based on the scale of the deposit and model block dimensions, 10 m long composites were deemed to be an appropriate length for estimating Mineral Resources.

 

14.4.4Geological Constraints – Iron Cap Deposit

 

Lithological, alteration, structural domains, and metal grade envelopes were constructed for the Sulphurets deposit by Seabridge and reviewed and accepted by the QP responsible for this section of this Report. Leapfrog software was used to generate preliminary wireframes that were further modified to account for various structural and lithological criteria.

 

After reviewing gold and copper distribution by lithology and alteration it was determined that independently constructed grade wireframes would be more appropriate for estimating block grades. Three structural domains were also used in the interpolation plan as hard boundaries.

 

14.4.5Variography – Iron Cap Deposit

 

Spatial continuity was investigated by generating both grade and indicator correlograms using Sage2001 software. Down-hole correlograms were used to establish nugget effects using 2 m long drill hole composites. Directional correlograms were generated using 10 m long drill hole composites in 30-degree vectors in plan and section. The autofit function in Sage2001 was used to model the 37 directional correlograms.

 

14.4.6Grade Estimation Parameters – Iron Cap Deposit

 

The Iron Cap deposit was modeled using MineSight® software, a widely recognized commercial mine engineering software package. The non-rotated model uses NAD83 UTM coordinates with 15 m x 15 m x 15 m blocks.

 

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Block gold, copper, silver and molybdenum grades were estimated by three primary methods: IDW, OK, and NN. The OK model was selected for resource declaration after comparing that model with the IDW model and NN models.

 

Independently constructed gold, copper, silver, and molybdenum wireframes provided the principal constraint in the grade estimation plan. In addition to the grade envelopes, three structural domain wireframes were developed to reflect different mineralized strike/dip regimes which allowed for specific search ellipses to be incorporated tor each structural block. The gradeshells were treated as soft contacts in the grade estimation plan where lower grade samples were allowed to inform the next highest grade shell but the higher grade samples were not allowed to inform adjacent lower grade zones.

 

A four-pass strategy was used for estimating block gold, silver, copper, and molybdenum grades for both the IDW and OK models. The first estimation pass used a 175 m x 175 m x 35 m search ellipse that required a minimum of five drill hole composites, a maximum of nine composites with no more than three composites per drill hole. That criteria resulted in blocks being estimated with at least two drill holes. All blocks estimated by the first pass were flagged and not overwritten by the subsequent estimation passes. The second pass used a 350 m x 350 m x 70 m search that required a minimum of five drill hole composites, a maximum of nine composites with no more than three composites per drill hole. The third estimation pass used a 500 m x 500 m x 100 m search ellipse and the same number of composites as pass one and two. A final cleanup pass used the same search ellipse as the third pass but only required a minimum of one drill hole.

 

NN models were generated for each metal at the same time the ordinary kriged grade was estimated. The NN models were used to access potential global biases in the grade estimates.

 

The number of composites and drill holes used to estimate each block along with distances to data (average and closest) were stored and used to help in the determination of Mineral Resource categories.

 

14.4.7Grade Model Verification – Iron Cap Deposit

 

Estimated block grades (gold, silver, copper, and molybdenum) were compared to drill hole composite grades in cross section and level plan views. Figure 14.20 and Figure 14.21 are northwest-southeast cross sections drawn through the Iron Cap block model (refer to Figure 10.5 for the location of Cross Section 12). Figure 14.22 and Figure 14.23 are block model level plans drawn at the 1200 m elevation through the Iron Cap block model showing estimated block grades and drill hole composite for gold and copper, respectively.

 

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Figure 14.20Iron Cap Cross Section 12 – Gold

 

 

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Figure 14.21Iron Cap Cross Section 12 – Copper

 

 

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Figure 14.22Iron Cap 1200 m Level Plan – Gold

 

 

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Figure 14.23Iron Cap 1,200 m Level Plan – Copper

 

 

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Grade swath plots were generated for rows (east-west), columns (north-south) and levels (elevations) through the block model comparing the OK and NN models at a zero cutoff grade. Figure 14.24 shows swath plots for gold (upper) and copper (lower) by elevation. These swath plots show Indicated and Inferred Resource grades.

 

Figure 14.24Iron Cap Gold-Copper Swath Plots by Elevation

 

 

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14.4.8Resource Classification – Iron Cap Deposit

 

Mineral Resources were assigned to the estimated blocks by constructing 3D solids for Indicated and Inferred Resources. These shapes were based on mineralized continuity as defined by exploration drill hole results. The average drill hole spacing for Indicated and Inferred blocks is approximately 75 m and 125 to 150 m, respectively. The drill hole spacing is wider at depth for Inferred Resources.

 

14.5Bulk Density

 

Bulk density values were assigned to each deposit based on an analysis of available specific gravity analyses that were performed on drill core during the core logging process. Bulk density determinations were routinely performed on core samples at a frequency of one sample per 100 m of drilling. Rare, apparent anomalously high or low determinations were excluded from the analysis. Most of bulk density data were assigned to each deposit by average densities by modeled lithology or alteration. Overburden and ice were assigned densities of 2.0 t/m3 and 0.9 t/m3 respectively. Bulk density values ranging between 2.7 t/m3 and 2.8 t/m3 were assigned to the four mineralized deposits.

 

14.6Resource Criteria

 

Based on calculated block NSR values, conceptual resource open pits and block cave underground shapes were generated for each model. Metal prices, costs (mining, processing and G&A), slope angles, and NSR cutoffs are summarized in Table 14.17.

 

Table 14.10Key Mineral Resource Parameters for Kerr, Sulphurets and Iron Cap

 

Parameter Units Value
Gold Price US$/oz 1,300
Copper Price US$/lb 3.00
Silver Price US$/oz 20.00
Molybdenum Price US$/lb 9.70
Conceptual Pit Mining Cost $/t 1.80
Conceptual Pit Processing + G&A Cost $t/process 9.00
Conceptual Pit Slope Angle Degrees 45
Conceptual Block Cave Mining Cost $/t 6.00 - 7.00
Conceptual Block Cave Processing + G&A Cost $t/process 9.00
Conceptual Pit Cutoff NSR Cdn $/t 9.00
Conceptual Block Cave Cutoff NSR Cdn $/t 16.00
Exchange US$/Cdn$ 0.83

 

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Table 14.11Key Mineral Resource Parameters for Mitchell and East Mitchell

 

Parameter Units Value
Gold Price US$/oz 1,300
Copper Price US$/lb 3.00
Silver Price US$/oz 20.0
Molybdenum Price US$/lb 9.7
Conceptual Pit Mining Cost $/t 2.2
Conceptual Pit Processing + G&A Cost $t/process 10.75 (Mitchell), 11.20
(East Mitchell)
Conceptual Pit Slope Angle Degrees 32-51 (Mitchell), 25-43
(East Mitchell)
Conceptual Pit Cutoff NSR (Mitchell)  Cdn $/t 10.75
Conceptual Pit Cutoff NSR (E. Mitchell) Cdn $/t 11.20
Exchange US$ / Cdn$ 0.83

 

14.7Summary of KSM Mineral Resources

 

Mineral Resources were established for the various KSM mineral deposits by:

 

constraining the estimates within geological models and grade shells,

 

constraining the estimates within either a conceptual open pit or conceptual block cave shapes,

 

applying an appropriate cutoff to the resource blocks

 

which established reasonable prospects for eventual economic extraction in accordance with the CIM Definition Standards (CIM, 2014).

 

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Table 14.12KSM Mineral Resources

 

Deposit Resource Category Tonnes (millions) Average Grades Contained Metal  
Gold (gpt) Copper (%) Silver (gpt) Moly (ppm) Gold ounces (millions) Copper pounds (millions) Silver ounces (millions) Moly pounds (millions)  
   
   
Kerr Measured  -  -  -  -  -  -  -  -  -    
Indicated Open Pit 351.0 0.22 0.41 1.1 4 2.5 3,172 12.4 3    
Indicated Block Cave 23.0 0.24 0.46 1.8 14 0.2 233 1.3 1    
Indicated Total 374.0 0.22 0.41 1.1 5 2.7 3,405 13.7 4    
Inferred Open Pit 78.0 0.27 0.21 1.3 6 0.7 361 3.3 1    
Inferred Block Cave 1,921.0 0.31 0.41 1.8 24 19.1 17,359 111.2 102    
Inferred Total 1,999.0 0.31 0.40 1.8 23 19.8 17,720 114.4 103    
Sulphurets Measured  -  -  -  -  -  -  -  -  -    
Indicated Open Pit 446.0 0.55 0.21 1 53 7.9 2,064 14.3 52    
Inferred Open Pit 223.0 0.44 0.13 1.3 30 3.2 639 9.3 15    
Mitchell Measured Open Pit 691.7 0.68 0.19 3.3 52 15.1 2,876 72.8 79    
Indicated Open Pit 1,667.0 0.48 0.14 2.8 66 25.9 5,120 149.2 241    
Inferred Open Pit 1,282.6 0.29 0.14 2.4 47 11.8 3,832 102.2 133    
East Mitchell Measured Open Pit 1,012.8 0.65 0.11 1.8 89 21.1 2,514 59.2 198    
Indicated Open Pit 746.2 0.42 0.08 1.7 79 10.1 1,390 41.8 130    
Inferred Open Pit 281.1 0.37 0.07 2.3 61 3.4 403 21.1 38    
Iron Cap Measured  -  -  -  -  -  -  -  -  -    
Indicated Block Cave 423.0 0.41 0.22 4.6 41 5.6 2,051 62.6 38    
Inferred Block Cave 1,899.0 0.45 0.3 2.6 30 27.5 12,556 158.7 126    

 

Notes:

 

1.The Mineral Resource estimate has been verified and endorsed Henry Kim P.Geo., an independent QP.

 

2.Resources are reported in accordance with the CIM Definition Standards (2014) and were estimated in accordance with the CIM Estimation of MRMR Best Practices Guidelines (2019).

 

3.Mineral Resources are reported inclusive of Mineral Reserves.

 

4.Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

 

5.The effective date for Mitchell and East Mitchell is March 31, 2021, and for Kerr, Sulphurets and Iron Cap is December 31, 2019.

 

6.Net Smelter Return (NSR) Cut offs are $10.75/t for Mitchell and $11.20/t for East Mitchell using the following assumptions: metal prices of US$1300/oz Au, US$3.00/lb Cu, US$20/oz Ag, and US$ 9.7/lb Mo at a currency exchange rate of 0.79 US$ per Cdn$; Copper concentrate terms are 96% payable Cu; 97.8% payable Au; 90% payable Ag. Offsite costs (smelting, refining, transport, and insurance) are $281 per tonne of concentrate; doré terms are $2/oz Au offsite costs (refining, transport and insurance), 99.8% Au payable, and 90% Ag payable; metallurgical recovery projections vary depending on metallurgical domain and metal grades and are based on metallurgical test work.

 

7.The Mineral Resource for the Mitchell and East Mitchell have been constrained by a “reasonable prospects of eventual economic extraction” pit using the following assumptions: metal prices of US$1820/oz Au, US$4.20/lb Cu, US$28/oz Ag, and US$ 13.5/lb Mo at a currency exchange rate of 0.83 US$ per 1.00 Cdn$; Pit slopes range between 32-51 degrees in the Mitchell area and 25-43 degrees in the East Mitchell area; $2.20/t mining costs; $10.75/t process + G&A costs for Mitchell; $11.20/t process + G&A costs for East Mitchell; offsite terms and metallurgical recoveries are the same as Note 6; Recoveries vary depending on metallurgical domain and metal grades and are based on metallurgical test work and see Section 13 for details.

 

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8.Net Smelter Return (NSR) Cut offs are $16/t for Iron Cap and Kerr block caves and $9/t for Sulphurets pit using the following assumptions: metal prices of US$1300/oz Au, US$3.00/lb Cu, US$20/oz Ag, and US$ 9.7/lb Mo at a currency exchange rate of 0.83 US$ per Cdn$; Copper concentrate terms are 96% payable Cu; 97.8% payable Au; 90% payable Ag. Offsite costs (smelting, refining, transport, and insurance) are $281 per tonne of concentrate; doré terms are $2/oz Au offsite costs (refining, transport and insurance), 99.8% Au payable, and 90% Ag payable; metallurgical recovery projections vary depending on metallurgical domain and metal grades and are based on metallurgical test work.

 

9.The Mineral Resource for Sulphurets deposit has been constrained by a pit shell using the following assumptions: metal prices of US$1300/oz Au, US$3.00/lb Cu, US$20/oz Ag, and US$ 9.7/lb Mo at a currency exchange rate of 0.83 US$ per Cdn$; Pit slopes for Sulphurets assumed 45 degrees; $1.80/t pit mining costs; $9/t process + G&A costs; offsite costs (smelting, refining, transport, and insurance) are $206 per tonne of concentrate; doré terms are $2/oz Au offsite costs (refining, transport and insurance), 99.8% Au payable, and 90% Ag payable; Recoveries vary depending on metallurgical domain and metal grades and are based on metallurgical test work and see Section 13 for details.

 

10.The Mineral Resource for Iron Cap deposit has been constrained by a conceptual block cave shape using the following assumptions: metal prices of US$1300/oz Au, US$3.00/lb Cu, US$20/oz Ag, and US$ 9.7/lb Mo at a currency exchange rate of 0.83 US$ per Cdn$; Mining cost of $6 - $7 per t; $9/t process + G&A costs; offsite costs (smelting, refining, transport, and insurance) are $206 per tonne of concentrate; doré terms are $2/oz Au offsite costs (refining, transport and insurance), 99.8% Au payable, and 90% Ag payable; Recoveries vary depending on metallurgical domain and metal grades and are based on metallurgical test work and see Section 13 for details.

 

11.The Mineral Resource for Kerr deposit has been constrained by “reasonable prospects of eventual economic extraction” open pit and block cave shapes using the following assumptions: metal prices of US$1300/oz Au, US$3.00/lb Cu, US$20/oz Ag, and US$ 9.7/lb Mo at a currency exchange rate of 0.83 US$ per Cdn$; Open pit mining cost of $1.80/t, underground mining cost of $6 - $7 per t; $9/t process + G&A costs; offsite costs (smelting, refining, transport, and insurance) are $206 per tonne of concentrate; doré terms are $2/oz Au offsite costs (refining, transport and insurance), 99.8% Au payable, and 90% Ag payable; Recoveries vary depending on metallurgical domain and metal grades and are based on metallurgical test work and see Section 13 for details.

 

12.“Moly” = “Molybdenum”

 

13.Numbers may not add due to rounding.

 

14.Unless noted otherwise, dollars reported herein are Canadian dollars.

 

14.8General Discussion

 

Mineral Resources could be materially impacted by the following risks:

 

Geological model may be more complex and may vary significantly in places from what has been modeled.

 

Block cave assumes a cutoff rather than a shutoff, and the dilution that will occur during mining will mean a portion of the Mineral Resources will be below the shutoff grade that will be applied during mining.

 

QP assumes the majority of the Inferred will be upgraded to at least Indicated with further drilling, but the portion that does not upgrade, and may be downgraded, would potentially reduce the size of the constraining mining shapes and reduce the mineral resource estimate.

 

Metallurgical recoveries are variable and dependant on grades. In particular, low grade Molybdenum can have lower grade portions of the deposits may not provide sufficient grade to be economically recoverable. There is a flexibility in the process plant to shut off the moly circuit when encountering low grade/low recoveries.

 

The QP responsible for this section of this Report is not aware of any specific environmental, permitting, legal, taxation, socio-economic, marketing, political or other relevant factors, other than what is identified in this Report, that could materially affect the Mineral Resource estimates that are the subject of this section.

 

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15.0Mineral Reserve Estimates

 

15.1Introduction

 

Mineral Reserves are based on modifying factors applied to Measured and Indicated Mineral Resources within a PFS mining study.

 

15.2Open Pit Reserve Parameters

 

The open pit minimum NSR cut-off grade is based on an estimated process operating cost of Cdn$11/t. Process operating costs include plant processing (including crushing and ore transport costs via tunnel to the mill), G&A, surface service, tailing construction, and water treatment costs. The NSR grade used for mine planning is described in Section 16.0. A variable cut-off grade strategy has been used; a higher cut-off grade of Cdn$25/t is used until the end of Year 5 to maximize the NPV. During this time material between Cdn$11/t and Cdn$25/t is wasted. The premium cut-off grade in the early years of the mine schedule assists in reducing the initial capital payback time.

 

15.3Mineral Reserves

 

Proven and Probable Mineral Reserves are summarized in Table 15.1 and match the production plan described in Section 16.0. The Qualified Person is not aware of any other risks, other than those identified in this Report, that could materially affect the Mineral Reserve estimates.

 

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Table 15.1Proven and Probable Reserves

 

  Ore
(Mt)
Diluted Grades Contained Metal
Au
(g/t)
Cu
(%)
Ag
(g/t)
Mo
(ppm)
Au
(Moz)
Cu
(Mlb)
Ag
(Moz)
Mo
(Mlb)
Proven Mitchell 483 0.74 0.20 3.3 49 11.5 2,161 51 53
East Mitchell 814 0.69 0.11 1.8 91 18.1 2,043 47 163
Sulphurets 0 0.00 0.00 0.0 0 0.0 0 0 0
Total Proven 1,297 0.71 0.15 2.4 75 29.6 4,203 98 215
Probable Mitchell 452 0.59 0.15 2.5 74 8.6 1,458 36 74
East Mitchell 392 0.46 0.09 1.7 84 5.8 784 21 73
Sulphurets 151 0.68 0.26 1.0 70 3.3 874 5 23
Total Probable 995 0.55 0.14 1.9 77 17.7 3,116 62 170
Proven + Probable Mitchell 935 0.67 0.18 2.9 61 20.1 3,619 87 126
East Mitchell 1,206 0.62 0.11 1.8 89 23.9 2,826 68 236
Sulphurets 151 0.68 0.26 1.0 70 3.3 874 5 23
Total Proven + Probable 2,292 0.64 0.14 2.2 76 47.3 7,320 160 385

Notes:

 

1.The Mineral Reserve estimates were reviewed by Jim Gray, P.Eng. (who is also the independent QP for these Mineral Reserve estimates), reported using the 2014 CIM Definition Standards and 2019 CIM Best Practices Guidelines, and have an effective date of May 26, 2022.
2.Mineral Reserves are based on the 2022 PFS all open pit Life of Mine plan.
3.Mineral Reserves are mined tonnes and grade, the reference point is the mill feed at the primary crusher and includes consideration for operational modifying factors.
4.Mineral Reserves are reported at NSR cut-off grades that vary between of Cdn$11/t and Cdn$25/t using the following assumptions: metal prices of US$1,300/oz Au, US$3.00/lb Cu, US$20/oz Ag, and US$9.7/lb Mo at a currency exchange rate of 0.79 US$ per Cdn$; Copper concentrate terms are 96% payable Cu; 97.8% payable Au; 90% payable Ag, molybdenum concentrate terms are 99% payable. Offsite costs (smelting, refining, transport, and insurance) are Cdn$281 per tonne of copper concentrate and Cdn$5,527 per tonne of molybdenum concentrate; doré terms are US$2/oz offsite costs (refining, transport and insurance), 99.8% Au payable, and 90% Ag payable; metallurgical recovery projections vary depending on metallurgical domain and metal grades and are based on metallurgical test work.
5.The NSR cut-off is varied from Cdn11/t to Cdn25/t and covers the estimated total process operating costs of Cdn$10/t for ore processing, G&A, surface service, tailings, and water treatment costs.
6.Mineral Reserves account for mining loss and dilution.
7.Mineral Reserves are a subset of the mineral resource
8.Numbers have been rounded as required by reporting guidelines.

 

15.4Factors that could affect the mineral reserve estimate

 

Mineral Reserves are based on the engineering and economic analysis described in Sections 16.0 to 22.0 of this Report. Changes in the following factors and assumptions could affect the Mineral Reserve estimate:

 

assumptions on weather and climate

 

effects of climate change

 

metal prices

 

interpretations of mineralization geometry and continuity of mineralization zones

 

interpolation assumptions

 

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geotechnical and hydrogeological assumptions

 

operating cost assumptions and price escalation

 

process plant and mining recoveries

 

ability to meet and maintain permitting and environmental license conditions

 

ability to maintain the social license to operate.

 

15.5Comments on Section 15.0

 

The current Mineral Reserve estimates are based on the most current knowledge, permit status, and engineering constraints. The QP is of the opinion that the Mineral Reserves have been estimated using industry best practices.

 

The Mineral Reserves as stated in this report do not include:

 

material within the designed pit that was sent to stockpile but not recovered to the mill in this plan,

 

mineralized material outside the designed pit where the pit did not mine out to the full extent of the economic pit limit. This material is potentially economic open pit or underground extraction for consideration in the future.

 

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16.0Mining Methods

 

16.1Introduction

 

A PFS-level production schedule, based on an annualized average 195,000 t/d mill feed rate, has been developed for the PFS based on open pit designs at Mitchell, East Mitchell, and Sulphurets deposits areas. The resulting mine life is approximately 33 years.

 

16.2Open Pit Mining Operations

 

16.2.1Introduction

 

The open pit mine planning work for this study is based on previous work including design criteria from the Application/EIS (Rescan 2013).

 

In addition to the geological information used for the block model, other data used for mine planning include the base economic parameters (metal prices, etc.), mining cost data derived from supplier estimates and data from other projects in the local area, recommended prefeasibility pit slope angles (PSAs), projected metallurgical recoveries, plant costs, and throughput rates.

 

16.2.2Mining Datum

 

The 2022 PFS design work is based on NAD83 coordinates. Historical drill hole information is based on various surveys with different sets of control that have been converted to NAD83. Topography is described in Section 12.1.6.

 

16.2.3Open Pit Mine Planning 3D Block Model

 

The block size is 25 m x 25 m x 15 m. The block heights represent a suitable bench height for large-scale mining shovels, and the block dimension are suitably sized for long-range planning.

 

Net Smelter Return

 

NSR per tonne (net of off-site concentrate treatment and smelter charges and including on-site mill recovery) is estimated for each block and is used as a cut-off item for break-even ore/waste selection.

 

NSR is estimated using net smelter price (NSP) and process recovery as shown in the equation below. The NSP is based on base case metal prices; US dollar exchange rate; and typical off-site losses, transportation, smelting, and refining charges. The terms of a final smelter schedule will be negotiated during the mine development. The major smelter terms used to estimate NSP are specified in Table 16.1, not including minor terms for deductions/losses, payables, price participation, etc.

 

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Where:

 

Cu = copper grade (%) from the CUIDW 3DBM item

Au = gold grade (g/t) from the AUIDW 3DBM item

Mo = molybdenum grade (ppm) from the MOIDW 3DBM item

Ag = silver grade (g/t) from the AGIDW 3DBM item

RecCu = copper recovery (%)

RecAu = gold recovery (%)

RecMo = molybdenum recovery (%)

RecAg = silver recovery (%)

NSPCu = net smelter price for copper (Cdn$/lb)

NSPAu = net smelter price for gold (Cdn$/g)

NSPMo = net smelter price for molybdenum (Cdn$/lb)

NSPAg = net smelter price for silver (Cdn$/g)

 

Table 16.1Major Smelter Terms Used in the NSR Calculation

 

  Amount Unit
Copper Concentrate
Smelting 100 US$/dmt
Au Refining 7.00 US$/oz
Ag Refining 0.50 US$/oz
Off-site Costs 281 Cdn$/wmt
Moly Concentrate
Roasting 2.00 US$/lb
Other Off-site Costs 5,527 Cdn$/wmt
Gold Dore
Au Refining + Transport 2.00 US$/oz

 

The metal prices and resultant NSPs used at this early stage of the study are shown by pit area in Table 16.2. Note that gold and silver NSP values are in Cdn$/gm.

 

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Table 16.2Metal Prices and Resultant NSP

 

  Metal Price
(US$)

NSP

(Cdn$)

Cu 3.00/lb 3.33/lb
Au 1,300/oz 48.6/gm
Ag 20.00/oz 0.674/gm
Mo 9.70/lb 6.80/lb
Exchange Rate (US$:Cdn$) 0.79  

 

Metallurgical recoveries used for the NSR calculation are based on test work conducted by G&T and evaluated by Tetra Tech and are described in Section 13.0.

 

Mining Loss and Dilution

 

The 2022 PFS involves a large gold-copper porphyry deposit and the orebody occurs relatively continuously within the cut-off grade shells. The pits will be mined with large shovels and trucks at an ore mining rate of 195,000 t/d. As is typical of large porphyries, blast hole assays will be used to determine the waste/ore boundaries for material designations on the pit bench for daily operations.

 

The Mineral Reserves used for scheduling are calculated from whole block grades in the 3DBM using detailed pit designs. The use of whole block grades for large size blocks means that dilution is already included in the block model. The use of autonomous trucks is assumed to eliminate loss along with the use of technology on the loading units (from misdirected loads).

 

16.2.4Pit Slope Design Angles

 

Overview

 

BGC has provided open pit slope design parameters for the three proposed open pits: Mitchell, Sulphurets, and East Mitchell. The design parameters are based on geotechnical site investigations, available local and regional geological data, and well-established geotechnical design methods.

 

BGC has identified geotechnical rock mass units associated with the primary rock and alteration types, based on the results of the site investigation and geological interpretations by Seabridge. Major geological structures (faults and foliation) have been included in the geotechnical slope stability analyses for each pit. Slope stability analyses were conducted using industry standard limit-equilibrium software, finite element analysis software, and in-house proprietary BGC tools.

 

BGC completed hydrogeological studies and simulations of pit dewatering/depressurization for the proposed Sulphurets and Mitchell pits. BGC interpreted hydrostratigraphic units, estimated hydraulic conductivity and storage parameter values, and formulated a conceptual hydrogeologic model for the study area. The conceptual model was used as the basis for developing a numerical hydrogeologic model. The calibrated numerical model was used to evaluate the effort required to depressurize the open pit slopes to satisfy geotechnical constraints identified in the open pit slope designs. Preliminary dewatering/depressurization plans, including the number of vertical wells, horizontal drains, and the extraction rates required to achieve sufficient depressurization of the rock mass, were developed to support the costing study. In addition, the need for a dewatering adit and associated drainage gallery was identified and simulated to achieve the depressurization targets of the upper north slope of the Mitchell pit. For the East Mitchell pit, preliminary dewatering/depressurization plans were developed based on the vertical well and horizontal drain spacing estimated for the other pits.

 

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Mitchell Pit Design

 

The proposed Mitchell pit will be located within a glacially modified valley and targets a mineral deposit located in the valley floor, resulting in 1,200 m high ultimate slopes. This scale of the Mitchell pit north and south slope heights will be similar to some currently operating, mature pits elsewhere in the Americas.

 

Multiple site investigation programs completed in the Mitchell Zone provide data for the Mitchell pit design work. Approximately 13,300 m of geotechnical drilling was completed, distributed over 31 core holes. Televiewer surveys were completed in 26 of these holes and oriented core measurements were collected in an additional 4 to provide geological discontinuity orientations for rock slope design. Packer testing was undertaken in 27 drill holes, and 28 vibrating wire piezometers were installed in 18 drill holes. Photogrammetric mapping of sections of the north and south valley walls was completed to provide additional data on the rock mass fabric of the study area.

 

Laboratory testing programs were completed, consisting of the following tests:

 

uniaxial compressive strength (70 tests)

 

triaxial compressive strength (13 tests)

 

Brazilian tensile strength (92 tests)

 

small scale direct shear testing (98 tests)

 

grain size and index testing (8 tests)

 

An appropriate quantity of data was collected to characterize the geological units of the study area and support PFS-level slope designs.

 

The structural geology of the Mitchell study area is defined by faults, foliation, and rock mass fabric (joints, etc.). In the Mitchell area the major faults are the Mitchell and Sulphurets thrust faults. The proposed Mitchell pit has been divided into four geotechnical domains, based on the different structural geology fabrics in the area; discontinuity sets and geotechnical units for each domain are identified for use in the slope designs. Design sectors are based on the anticipated main orientations of the proposed pit walls, as determined from previous pit optimization studies.

 

Recommended inter-ramp slope angles vary from 36° to 53° based on wall orientation, overall wall height, geotechnical domain, and controls on slope stability. Inter-ramp slope heights are limited to 150 m, after which a geotechnical berm (or ramp) with a minimum width of 25 m is required. The inter-ramp height limits and geotechnical berms provide flexibility in the mine plan to mitigate potential slope instability; access for slope monitoring installations; and working space for in-pit wells, drains, and other water management infrastructure. All final pit slopes are assumed to be excavated using controlled blasting. Depressurization of the proposed pit slopes requires a combination of vertical wells, horizontal drains, and a dewatering adit with drainage galleries.

 

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Sulphurets Pit Design

 

The proposed Sulphurets pit will be located on a glacially modified ridge between the Mitchell and Sulphurets valleys. The proposed mine plan would result in ultimate pit slopes with maximum heights of approximately 630 m, and a footprint of approximately 2 km x 1 km, with the long axis of the pit trending parallel to the strike of the STF.

 

A site investigation program including geotechnical drilling and hydrogeological testing was completed in 2010. Data from five geotechnical drill holes (consisting of approximately 1,950 m of drilling) was used to divide the Sulphurets Zone into three geotechnical domains: the hanging wall of the STF, the footwall of the STF, and an altered (crackled) zone associated with and defined by the STF. Additional joint and bedding sets have also been identified.

 

Laboratory testing of core samples from the completed geotechnical drilling included:

 

uniaxial compressive strength (13 tests)

 

Brazilian tensile strength (20 tests)

 

small scale direct shear tests of natural discontinuities (5 tests)

 

index testing of discontinuity infilling material (3 tests).

 

The rocks of the Sulphurets Zone are typically moderately strong when weathered and strong when fresh. The RQD of the rocks of the Sulphurets Zone varies from fair to good, generally increasing in quality with depth below surface or distance from the STF.

 

The slope designs assume final walls will be excavated using controlled blasting, consistent with the approach proposed for the Mitchell pit. The recommended inter-ramp slope angles vary from 36° to 50° based on wall orientation, overall wall height, rock mass quality, and structural controls on slope stability. Inter-ramp slope heights are limited to 150 m after which a geotechnical berm (or ramp) with a minimum width of 20 m is required. Depressurization of the pit slopes is required and should be achievable with a combination of vertical wells and horizontal drains.

 

East Mitchell Pit Design

 

The proposed East Mitchell open pit is located on the south side of the Mitchell Valley to the east of the Mitchell pit and the south of the Mitchell Glacier. The proposed mine plan will result in ultimate pit slopes approximately 1,100 m high, with a proposed pit footprint of approximately 2 km x 2 km.

 

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Several site investigation programs in the East Mitchell Zone were completed from 2009-2021. These programs include 21 geotechnical drill holes, consisting of approximately 7,150 m of drilling. Collected geotechnical data were used to divide the East Mitchell into six geotechnical domains based on the different structural geology fabrics in the area; discontinuity sets and geotechnical units for each domain are identified for use in the slope designs. The structural geology of the East Mitchell Zone includes discontinuity sets parallel to the MTF, STF, and Brucejack Fault, as well as bedding and foliation.

 

Laboratory testing of core samples from the geotechnical drilling included:

 

uniaxial compressive strength (62 tests)

 

triaxial compressive strength (1 test)

 

Brazilian tensile strength (101 tests)

 

small scale direct shear tests of natural discontinuities (123 tests)

 

index testing of discontinuity infilling material (19 tests).

 

The rocks comprising the proposed pit slopes are generally medium strong to strong with a good rock mass rating.

 

The slope designs assume that final walls will be excavated using controlled blasting. The recommended inter-ramp slope angles vary from 27° to 45° based on overall wall height, wall azimuth, rock mass quality, and geological structures. Inter-ramp slope heights are limited to 150 m after which a geotechnical berm (or ramp) with a minimum width of 25 m is required. Depressurization of the pit slopes is required and should be achievable with a combination of vertical wells and horizontal drains.

 

Slope Design Implementation

 

Achieving the slope designs will require depressurization of the pit walls through the use of vertical wells and horizontal drains. Geological structures may affect bench and inter-ramp scale slope stability, and therefore, depressurization of these structures will be required.

 

Based on groundwater modelling results, approximately 76 in-pit wells will be required over the LOM for the Mitchell pit. The total drilling length for the vertical wells is estimated to be approximately 15,200 m. In addition, a 3.7 km adit and drainage gallery will be required for the Mitchell pit north wall, and approximately 764 km of horizontal drains will be required to aid in depressurization of the pit slopes over the mine life.

 

For the East Mitchell pit, approximately 69 vertical wells with a total drilling length of 13,800 m will be required throughout the life of the pit. In addition, it is estimated that approximately 277 km of horizontal drains will be required to aid in depressurization of the pit slopes over the life of the pit.

 

For the Sulphurets pit, approximately 10 vertical wells with a total drilling length of 2,000 m will be required throughout the life of the pit. In addition, it is estimated that approximately 65 km of horizontal drains will be required to aid in depressurization of the pit slopes over the life of the pit.

 

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Further characterization of the hydraulic properties of the bedrock at the feasibility stage and dewatering and depressurization response must be monitored throughout mining operations to determine if targets are being met.

 

Monitoring of pit slope displacements at various scales will be required. Inter-ramp and overall scale slopes should be monitored for deformations. The slope deformation monitoring system designed for the Mitchell and East Mitchell pits will meet or exceed the capability of those systems currently in operation at other large open pits elsewhere. The monitoring system should include multiple robotic total stations and survey prisms, mobile slope stability radar units, slope inclinometers, piezometers, and extensometers. The system would be computerized and use radio telemetry or a similar technology to provide real-time data to on-site geotechnical and mining staff. Similar monitoring systems would also be required for the Sulphurets pit; the requirements of those systems would be scaled according to the proposed wall heights for those pits.

 

It will be important to manage geological hazards during mining operations. Additional engineered structures adjacent to the pit, or modifications to the pit slope geometry, may be required to mitigate the risk of snow avalanches. In addition, the area has been recently de-glaciated and large-scale slope deformation features have been identified in the Mitchell and Sulphurets valleys.

 

Nine large landslides in the study area will require management during construction and operations. Conceptual management plans detailing monitoring and mitigation measures have been prepared to mitigate risk associated with existing landslides. Of particular note with respect to the open pits is the Snowfield Landslide situated on the south slope of the Mitchell Valley, which will be mined out by the East Mitchell pit.

 

The overall landslide management plan uses a risk-based approach to determine the level of monitoring required for each landslide. The management plan for the Snowfield Landslide is comprehensive due to its proximity to the Mitchell open pit. The plan includes surface and subsurface deformation monitoring, surface water management, pumping wells, and a depressurization adit.

 

16.2.5Economic Pit Limits, Pit Designs

 

pit optimization method

 

The economic pit limit is selected after evaluating LG pit cases.

 

The assessment is carried out by generating sets of LG pit shells by varying revenue assumptions to test the deposit’s geometric/topographic and pit slope sensitivity.

 

LG Pit Assumptions

 

Inputs to the updated LG pit limit assessment shown in Table 16.3 are based on the previous PFS studies (considering cost escalation) as a starting point for the 2022 design work.

 

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Table 16.3LG Pit Limit Primary Assumptions

 

Assumption Value
Mining Cost Cdn$2.20/t
Process, G&A, Site Services, Water Treatment - Mitchell Cdn$10.75/t
Process, G&A, Site Services, Water Treatment – East Mitchell Cdn$11.20/t
Process Recoveries See Section 17.0
Pit Slope Angle Variable See Section 16.2.4
Metal Prices

See
Table 16.2

 

 

LG pits are generated by varying prices in the range from 40% to 150% of the base NSR.

 

LG Economic Pit Limits

 

The ultimate OP limits were selected based on a combined maximum ore constraint of 2.3B tonnes (TMF capacity) and keeping the Mitchell pit north wall crest from advancing onto the upper elevation steeper natural slope angles. The sensitivity of the LG economic pit limit to net revenues for Mitchell and east Mitchell limit is illustrated in Figures 16.1 & 16.2 where 100% is from the base case NSP parameters above.

 

The selected open pit limits are summarized below:

 

Mitchell – 60% Case

 

East Mitchell – 80% Case

 

Sulphurets – 2016 phase 1 pit

 

The selected cases for Mitchell and East Mitchell are well within the economic limits for this study. The Sulphurets selection is the first phase of the previous 2 phase ultimate pit design is the higher grade/lower strip ratio, and is well within the Sulphurets economic pit limit.

 

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Figure 16.1Mitchell Sensitivity of Ore Tonnes to Pit Size

 

 

Source: MMTS, 2022

 

Figure 16.2East Mitchell Sensitivity of Ore Tonnes to Pit Size

 

 

Source: MMTS, 2022

 

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A plan view and north-south section views of the LG pits for the open pit mining areas are shown in Figure 16.3 through Figure 16.5.

 

Figure 16.3Plan View of the KSM LG Pit Limits

 

 

Source: MMTS, 2022

 

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Figure 16.4Mitchell Economic Pit Limit – North-South Section at East 422950, Viewed from the East

 

 

Source: MMTS, 2022

Note: UNSR2 is NSR values in Cdn$ per tonne

 

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Figure 16.5East Mitchell Economic Pit Limit – North-South Section at East 424725, Viewed from the East

 

 

Source: MMTS, 2022

Note: UNSR2 is NSR values in Cdn$

 

16.2.6Detailed Pit Designs

 

PFS-level pit designs demonstrate the viability of accessing and open pit mining the Measured and Indicated Mineral Resources at the KSM site. Pit designs use the selected LG pit limits as guides as well as geotechnical parameters, suitable road widths, and minimum mining widths based on efficient operation for the size of mining equipment chosen for the 2022 PFS.

 

Haul Road Widths

 

Haul road widths are designed to provide safe, efficient haulage and to comply with the BC Mines Regulations’ minimum width specifications and safe operating practice. Haul road widths include allowance for standard double lane hauling, as well as an allowance for a single trolley lane. All roads have a maximum grade of 8% suitable for mines with winter conditions. A sample cross section of a haul road is shown in Figure 16.6 below.

 

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Figure 16.6Double Lane Haul Road with Trolley Lane

 

 

(MMTS, 2022)

 

Design Standards

 

Detailed design parameters for pits and RSFs are provided by BGC and KCB, respectively, according to their geotechnical testing and evaluations (Sections 16.2.4 and 18.2).

 

Minimum Mining Width

 

A minimum mining width between pit phases is prescribed to maintain a suitable platform for efficient mining operations. This is established based on equipment size and operating characteristics. For this study, the minimum mining width generally conforms to 50 m, which provides sufficient room for 2-sided truck loading.

 

Variable Berm Width

 

Pit designs for KSM are designed honouring overall PSAs, a nominal bench face angle (60° to 70°), and variable safety berm widths with a minimum 8 m width (BC mining regulation). Due to the low overall PSAs and double benching between berms, berm widths are generally greater than 15 m. Where haul roads intersect designed safety benches, the haul road width is counted towards the safety berm width for the purpose of calculating the maximum overall PSA.

 

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Bench Height

 

The KSM pit designs are based on the digging reach of the large shovels (15 m operating bench) with double benching between high wall berms; therefore, the berms are separated vertically by 30 m. Single benching will be employed, if required, to maximize ore recovery and maintain the safety berm sequence as warranted.

 

LG Phase Selection

 

The LG selected pit cases discussed previously are used to evaluate alternatives for determining the economic pit limit and the optimal push-backs or phases before commencing detailed design work.

 

There are smaller pit shells within the economic pit limits that have higher economic margins, due to their lower strip ratios or better grades than the full economic pit limit. Mining these pits as phases from higher to lower economic margins maximizes revenue and minimizes mining cost early in the mining operations.

 

The description of the detailed pit designs and phases in this section uses the following naming conventions:

 

The letters M, EM, and S signify Mitchell, East Mitchell, and Sulphurets, respectively.

 

The digit signifies the pit phase number.

 

A suffix of ‘i’ indicates that the reserve tonnage for the phase is incremental from the previous phase. If there is no ‘i’ specified, it is cumulative within the pit, up to the phase indicated.

 

Mitchell Pits

 

Mitchell pit has four incremental phases. Pit phase M1 is the quarry designed to provide pre-production start-up and construction material. Phases M2i and M3i mine to the south, with M3i reaching the ultimate pit limit in the south. The final pit phase, M4i is a high strip ratio pushback, that mines to the ultimate pit limit in the north.

 

The Mitchell pit phases have been designed to mine vertically through the Snowfield Landslide on the southeast side of the pit and not undermine it.

 

A plan view of the Mitchell pit phases is shown in Figure 16.7.

 

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Figure 16.7Plan View of Mitchell Pit Phases

 

 

Source: MMTS (2022)

 

See Figure 16.3 for location within the claim boundary

 

East Mitchell Pits

 

East Mitchell pit has three incremental phases. Each progressive phase pushes back to the south.

 

A plan view of the East Mitchell pit phases is shown in Figure 16.8.

 

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Figure 16.8Plan View of East Mitchell Pit Phases

 

 

Source: MMTS (2022)

 

See Figure 16.3 for location within the claim boundary

 

Sulphurets Pit

 

The Sulphurets deposit is mined as a single phase.

 

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Figure 16-9Plan View of Sulphurets Pit

 

 

Source: MMTS (2022)

 

See Figure 16.3 for location within the claim boundary

 

16.2.7Open Pit Mine Plan

 

LOM Open Pit Production Schedule

 

The open pit mine production schedule is developed with MinePlan Strategic Optimizer (MPSO), a comprehensive long-range schedule optimization tool for open pit mines used to produce a LOM schedule that increases the NPV of the PFS.

 

In the open pit mine schedule, “Time 0” refers to the mill start date; full mill feed production capacity is expected at the start of Year 2. The production schedule specifies:

 

pioneering: Year -6, -5, -4, -3

 

pre-production: Year -2, -1

 

first year of operations: Year 1

 

Open Pit Mine Load and Haul Fleet Selection

 

The mine load and haul fleet are selected prior to production scheduling. Previous studies and similar projects in the area have shown that the lowest cost per tonne fleet of cable shovels and haul trucks that are currently being used for large hard rock open pit mines are the 100 t bucket class shovel matched with the 360 t truck. This size of equipment is proven in operating mines around the world. The decision has also been made to include overhead electric catenaries (trolley assist) to the haul roads for use by the haul trucks to speed up the loaded trucks and reduce diesel consumption.

 

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Diesel hydraulic shovels (85 t bucket class) are used when a more mobile loading unit is needed. Suitable drill sizes (311 mm hole size) are chosen to match this size of truck/shovel fleet. The following performance and costs are estimated based on the use of these large-scale mining equipment.

 

Productivities of the selected equipment include shovel loading times and truck haul cycle estimates for multiple pit-to-destination combinations.

 

Cycle time and fuel burn reductions due to utilization of trolley systems are incorporated into the haul truck operating hour calculations. An overall trolley utilization factor is included to represent the availability of the trolley system (when the truck is on trolley) as well as the availability of a trolley line on the haul network (accounts for lag between road being opened up and the installation or extension of the trolley line) Trolley lines are assumed on ramps that are greater than 4% and a minimum trolley length of 500 m is assumed.

 

Autonomous drills and trucks are used during regular mining operations. Tele-remote operated shovels are also assumed. Autonomous and remotely operated units have a lower mechanical availability (since there is more time required to maintain the technology on board) but a higher operator efficiency and increased operating hours per day (no breaks or shift-change delays) Standard delays such as blasting, fueling, clean-up, long moves, etc., are still accounted for with autonomous units.

 

Schedule Criteria

 

In order to optimize the PFS NPV, grade bins have been specified (based on NSR block values); the MPSO develops a cut-off grade strategy to increase the PFS NPV. This increases mill head grades, and therefore revenues, early in the production schedule. Pit/phase precedencies are specified based on the logical progression of phase geometry (no undermining) but also determined by the timing of water diversions, bench access issues, backfill opportunities, and RSF phase sequencing.

 

The primary program objective in each period is to maximize the NPV. The MPSO NPV calculation is guided by estimated operating and capital costs, process recoveries, and metal prices. Key production schedule assumptions are shown in Table 16.4.

 

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Table 16.4Production Schedule Assumptions

 

Assumption Value
Operating Days Per Year 355
Hours Per Day 21
Daily Mill Throughput (Years 1 & 2) 130,000 t/d
Daily Mill Throughput (After Year 2) 195,000 t/d
Haul Truck Speed Limit 50 km/h
Haul Truck Operator Efficiency - autonomous 100%
Haul Truck Operating Efficiency - autonomous 90%
Dump and Manoeuvre Time 1.5 min
Shovel Loading Time 30 s/pass
Shovel Operator Efficiency 95%
Shovel Operating Efficiency 93.3%

 

Allowance has been made for the severe snowstorms or when poor visibility requires the mine to partially or completely shut down. The results from previous conditional simulations with respect to weather and storms and the effect on open pit mine traffic has been applied to quantify some of the lost operating time. Allowance for fog and other conditions have also been deducted. The cumulative lost production hours are applied as lost days in the schedule.

 

Cut-off Grade Optimization

 

The pit phase designs and sequencing are typically from higher grades to lower, to mine the higher mill feed grades early in the schedule and thereby increase the PFS revenues in the earlier years. This can be further enhanced by stockpiling low- and mid-grade. The stockpiled material is then milled at the end of the production schedule. However, stockpiling also results in increased mining cost per t milled. Additionally, oxidation can cause significant metallurgical recovery loss in the stockpile. At some point, the cost of mining more material and the recovery loss will exceed the incremental benefit from the higher grade milled. A variable cut-off grade strategy has been applied for the KSM production schedule to improve the 2022 PFS economics by mining the best grade ore first, varying the stockpiling, and smoothing out the haul fleet to optimize the NPV.

 

Rock Storage Facilities

 

The RSF is located as close to the mining areas as possible. Mitchell, East Mitchell and Sulphurets waste is placed in Mitchell RSFs. East Mitchell Phase 3 waste is placed into the mined out Mitchell pit.

 

Further details on the RSF design are available in Section 18.0.

 

Construction Methods

 

Several different construction methods will be used for waste placement: top-down, bottom-up, and wraparounds. Top-down platform heights are limited to a maximum height of 300 m. Bottom-up lifts are 30 m to 50 m high, or less if geotechnically required. Wraparounds are built onto the face of an existing RSF, creating a series of terraces to facilitate intermediate haul roads and lower the overall slope angle of high dumps, which will reduce the end of mine closure costs.

 

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Foundation Preparation

 

Design work for RSF foundation preparation will be performed as required at the feasibility-level design stage. Prior to mine development, soil will be salvaged from the footprint area where soil is suitable for reclamation purposes. Soils will generally not be salvaged on slopes steeper than 26° due to practical limitations on equipment access and operator safety. Soils salvaged from the RSF footprints will be stockpiled in the Ted Morris Valley.

 

The waste in the valley bottoms will be placed in low height lifts to confine and consolidate weaker foundation material. If required, loose tills and clays will be removed. Once the foundation is prepared, the basal drain is placed on top at the required lift height. The basal drain requirements are described in Rescan (2013).

 

RSF Monitoring and Planning

 

The long-term operation of the RSFs will be similar to that of the large, steep-terrain RSFs that have been in operation for many years in southeast BC Rocky Mountain coal mines. These operations involve high-relief RSF phases with clear dumping in single lifts of up to 100 m. Clear dumping is a technique whereby truck loads are dumped directly over the crest of the dump face; the load is not dumped short and then pushed over the edge. The clear dumping technique better maintains a stable dump platform at the crest but requires well-established monitoring and operating practices. Foundation preparation also needs to be assured.

 

RSF Access Roads

 

Pioneering access to each pit and subsequent phases use roads with a maximum 15% grade; these are constructed using balanced cut and fill wherever possible. Pioneering roads are 10 m wide and enable major mining equipment to reach the top benches of each pit phase and start mining. These are built for pit access and not for hauling. After the pioneering road is established to the top benches of each pit phase, bench waste from the upper portions of each pit phase is used to fill full-width haul roads at a maximum gradient of 8% at the appropriate double lane width, to connect with permanent surface roads and high wall roads in the long-term road network. This road network connects the mining areas with the primary crusher and stockpile areas for ore and the RSF areas for waste.

 

As described earlier, the terraced RSFs on the south side of the Mitchell Valley provide level access to the south Mitchell Valley RSF platforms.

 

Final RSF Configuration

 

The RSF for the PFS will have overall slope angles of 26° to 30°. The final post-closure configuration will be adapted in accordance with the closure plan as identified within Rescan (2013) and described in Section 20.6. Costs for this work are included during the later years of the operation, when ancillary equipment then becomes available for other duties.

 

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Low-grade ROM Stockpile

 

Lower-grade ore is stockpiled throughout the mining schedule to follow the cut-off grade strategy, and then is reclaimed in later years. This increases the grade of the ore feed to the plant in the early years of the schedule, but also is used to even out the waste mining requirements as required during periods of high pre-stripping for some of the pit phases. The low-grade ore stockpile is placed to the west of the Mitchell OPC. Provision has been made for an HDPE pipeline diversion around the surface of the stockpile, which can be moved as required.

 

Open Pit Mine Pre-production Detail

 

Development and pre-production activities include the following:

 

Expose sufficient ore for start-up.

 

Establish mining areas that will support the equipment required to achieve ore production and annual mill feed requirements on a sustainable basis.

 

Provide material required for construction in the mine area.

 

In early development, the mining fleet will excavate material from M1 to provide construction fill for the Mitchell OPC, Water Storage Dam, and the basal drain.

 

During pre-production, Mitchell pit phase M2 is mined to 1,080 m on the south side and 1,110 m on the north side. During the first half of Year -2 the south side of M2 is mined using tele-remote dozers to push material from south to north, spilling down into the Mitchell valley below. This creates rehandle on lower benches but allows a quicker vertical advance rate since double-bench blasting is used and only pioneering roads are required (rather than full size haul roads) as no waste is being hauled externally. Once the dozer push distance reaches ~50 m (around the 1,380 m elevation), the productivity starts to drop off and tele-remote shovels loading autonomous trucks are used for the portions of the in-situ material along the highwall. The trucks will short-haul this material and dump it over the outside edge. This also creates rehandle but it allows for easier benching of the pre-strip down to first ore with fewer trucks during pre-production and start-up. Shovel/trucks are used to mine anywhere between 20-35% of the material on each bench elevation from 1,380 m down to 1,080 m on the south side. The north side is done with dozers only. In total there are 533M tonnes of material moved in M2 during pre-production, 305 million t in-situ, and 228M tonnes rehandle. Tele-remote and autonomous equipment allows for higher productivities and decreases personnel exposure.

 

Development and pre-production must be completed by the end of Year -1. This will expose sufficient ore for commissioning at 130,000 t/d of mill feed after that. The mill feed will increase to 195,000 t/d at the beginning of Year 3 until the end of LOM. The mine layout at the end of pre-production is shown in Figure 16.10.

 

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16.2.8Open Pit Production

 

Year 1 to 33 – Open Pit Mining

 

The following is a summary of mining activity in Years 1 to 5:

 

Mining in Year 1 to 5 focuses on delivering the grade of ore required to payback initial capital

 

All waste material is placed in the Mitchell RSF

 

Ore is hauled directly to the Mitchell OPC

 

An ROM ore stockpile is built in the area to west of the Mitchell OPC

 

The Mitchell RSF is built in lifts at an overall slope of 2:1 with an access road in the final face

 

The following is a summary of mining activity in Years 6 to 33:

 

By Year 6, M2 is mined out and mining continues in the upper parts of EM1, EM2 and Sulphurets

 

Sulphurets is mined out by Year 12

 

Mining continues in EM1 and EM2, with both phases finishing in Year 18

 

M3 begins mining in Year 15 and is completed in Year 21

 

M4 begins mining in Year 17 and is completed in Year 23

 

EM3 begins mining as M4 is finishing. Once M4 is finished, the waste from EM3 is placed into the Mitchell pit as backfill. EM3 is mined out in Year 33

 

Stockpile material is reclaimed to supplement mill feed during periods where mining is limited by periodic large volumes of waste pre-stripping or to reduce vertical advance rate as required.

 

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Figure 16.10End of Pre-production (Year -1)

 

 

Source: MMTS (2022)

 

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Figure 16.11End of Open Pit Life of Mine

 

 

Source: MMTS (2022)

 

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16.2.9Open Pit Mine Operations

 

KSM open pit mining operations will employ bulk mining methods and large capacity equipment. Services and support are well established in BC and in the local area for this kind of operation.

 

The use of large-scale equipment will minimize unit operating costs, reduce the on-site labour requirement of a remote site, and dilute the fixed overhead costs for the open pit operations.

 

Organization

 

Mine operations is organized into three areas: direct mining, mine maintenance, and general mine expense (GME).

 

In this study, direct mining and mine maintenance are planned as an owner-operated fleet with the equipment ownership and labour being directly under mine operations. The alternative of to contracting out some of the direct mining activities under typical mine stripping contracts, and maintenance and repair contracts (MARC) has not been included in this study. The mine will employ the blasting crew but the supply and onsite manufacturing of blasting materials will be contracted out. All infrastructure required for the blasting supply contractor will be provided by the operations.

 

Direct Mining Activities – Open Pit

 

The direct mining area accounts for the drilling, blasting, loading, hauling, and pit maintenance activities in the mine.

 

Drilling

 

Blasthole drills will be fitted with GPS navigation and drill control systems to optimize drilling. Blasthole drills will be operated autonomously.

 

Electric rotary drills (311 mm bit size) will be used for production drilling, both in ore and waste. Diesel hydraulic percussive drills (150 mm bit size) will be used for controlled blasting techniques on high wall rows, pioneering drilling during pre-production, and development of initial upper benches.

 

blasting

 

A contract explosives supplier will provide blasting materials and technology. Due to the remote nature of the operation, an explosives manufacturing plant will be built on site when emulsion is required. For this study, the owner provides a serviced site and all facilities to the explosives contractor who manufactures and delivers the prescribed explosives to the blast holes and supplies all blasting accessories.

 

Blasting will be done with a 35/65 emulsion/ammonium nitrate-fuel oil (ANFO) mix explosive. Higher use of straight ANFO with borehole liners to keep the ANFO dry, can be investigated in future studies to reduce blasting costs.

 

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Blasting accessories will be stored in magazines adjacent to the mining areas suitably located to meet federal and provincial regulations and to avoid potential geohazards.

 

Loading of the explosives will be done with bulk explosives loading trucks provided by the explosives supplier. The trucks should be equipped with GPS guidance and should be able to receive automatic loading instructions for each hole from the engineering office. The GPS guidance will be a necessity to be compatible with stakeless drilling.

 

A smaller “goat” truck is needed for small development areas as well as for squaring-off blast patterns when the mine roads have been closed due to excessive snow fall. “Goat” trucks are similar to a logging skidder with high manoeuvrability and are a specific adaptation for open pit operations in mountainous and high snow fall areas.

 

Blast holes will be stemmed to avoid fly-rock and excessive air blasts. Crushed rock will be provided for stemming material and will be dumped adjacent to the blast pattern. A loader with a side dump bucket is included in the mine fleet to tram and dump the crush into the hole.

 

The blasting crew will comprise mine employees and will be on day shift only. The blasting crew will coordinate drilling and blasting activities to ensure a minimum of two weeks of broken material inventory is maintained for each shovel. Blasters will require hand-held GPS to identify holes for pattern tie-in as blast patterns will not be staked. A detonation system will be used that consists of electric cap initiation, detonating cord, surface delay connectors, non-electric single-delay caps, and boosters.

 

Loading

 

Ore and waste will be defined in the blasted muck pile by the OCS. A fleet management system will assist in optimizing deployment and utilization of the mine fleet

 

Three 85-t dipper diesel hydraulic shovels and three 100-t dipper electric cable shovels have been selected as the primary digging units. The diesel hydraulic shovels are selected for flexibility and mobility in accessing the narrow top pit benches.

 

Minimum bench widths of 50 m are designed to ensure sufficient operating for double-sided loading of trucks at the shovels. Where single-sided loading will be necessary and reduced productivity for the shovel will be encountered, such as the upper benches of the pit phases ancillary equipment will be deployed to prepare the digging areas for higher shovel productivity. This can entail dozing small benches down slope to the next bench, trap dozing, and other dozing activities.

 

Hauling

 

Ore and waste will be hauled by 360 t off-highway haul trucks. Haul productivities have been estimated from each bench to designated dumping points as the dump progresses. The 360 t trucks will be configured for overhead trolley assist to reduce fuel use and improve haul speeds. Catenary lines will be constructed on haul road segments that service significant haulage requirements.

 

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Pit Maintenance

 

Pit maintenance services include haul road maintenance, open pit mine dewatering, transporting operating supplies, relocating equipment, and snow removal.

 

Haul road maintenance is paramount to low haulage costs; dozer and grader hours have been allocated to maintain the haul road network throughout the LOM production schedule.

 

A fleet of ancillary service vehicles are allocated to install and service the in-pit sump pumps and the high wall horizontal drains. This includes connecting these pumps to the pit dewatering pipeline system. This crew will also service and supply mobile light plants.

 

A fleet of service equipment is allocated for summer season construction and will be used in winter for snow clearing and spreading crushed rock for traction control. This includes scrapers and loaders. The snow fleet will be manned by mine operations staff in normal winter conditions with operators taken from reduced activities such as dust control and summer field programs. During severe storms, personnel to operate the standby snow fleet will be drawn from truck and shovel operations as the long-haul fleets shut down. This will ensure priority fleets remain operating.

 

A rock crusher for road grading material is included.

 

Open Pit General Mine Expense Area

 

The GME area accounts for the supervision, safety, environment, and training for the direct mining activities as well as technical support from mine engineering and geology functions. Open pit mine operation supervision will extend down to the shift foreman level and trainers.

 

GME costs also include engineering consulting expenses on an ongoing basis for specialty items, such as geotechnical and geo-hydrology expertise, and third-party reviews in the open pit mine area.

 

16.2.10Mine Closure and Reclamation

 

Details on mine closure and reclamation are available in Section 20.6

 

16.2.11Open Pit Mine Equipment Parameters

 

Mining equipment descriptions in this section provide general specifications so that dimensions and capacities can be determined from vendor specification documents.

 

Major Equipment

 

The production requirements for the major mining equipment over the LOM are summarized in Table 16.5. The current production schedule requires a maximum haulage fleet of 55 trucks over the LOM.

 

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Table 16.5Major Equipment Requirements

 

  Pre-
production
Year
5
Maximum
Primary Drill – 311 mm Electric Drill 6 9 9
High Wall Drill – 150 mm Diesel Hydraulic Drill 3 5 5
Primary Shovel – 40 m3 Diesel Hydraulic Shovel 2 3 3
Primary Shovel – 56 m3 Electric Cable Shovel 0 3 3
Construction FEL – 12 m3 1 2 2
Haul Truck – 360 t 6 42 55
Construction Haul Truck – 90 t 2 3 3

 

blasting

 

Blasting assumptions are summarized in Table 16.6. These parameters are typical for other mines in western Canada and will be re-evaluated in the future with a detailed blasting study, using site-specific rock strength parameters.

 

Table 16.6Blasting Assumptions

 

Blasting Pattern – Ore and Waste Specifications
Spacing 8.15 m
Burden 8.15 m
Hole Size 12¼ inches
311 mm
Explosive In-Hole Density 1.15 g/cc
Explosive Average Downhole Loading 87.4 kg/m
Bench Height 15 m
Collar 6 m
Loaded Column 11 m
Sub-drill 2 m
Charge per Hole 961 kg/hole
Rock SG 2.77 t/m3
Yield per Hole 2,760 t/hole
Powder Factor 0.35 kg/t

 

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Drilling Equipment

 

Production drilling assumptions are listed in Table 16.7.

 

Table 16.7Open Pit Production Drilling Assumptions

 

Production Drill –
Mineralized
Material & Waste
Electric
Rotary
Diesel
Rotary
Bench Height 15 m 15 m
Subgrade 2.0 m 2.0 m
Hole Size 311 mm 311 mm
Penetration Rate 40.0 m/h 40.0 m/h
Hole Depth 18 m 18 m
Over Drill 1.0m 1.0 m
Setup Time 2.0 min 2.0 min
Drill Time 27.0 min 27.0 min
Move Time 2.0 min 2.0 min
Total Cycle Time 31.0 min 31.0 min
Holes per Hour 1.94 1.94
Re-drills 6% 6%

 

A 150 mm diesel percussive drill is also specified for controlled blasting techniques on high wall rows in all pit phases, pioneering drilling during pre-production, and development of initial upper benches.

 

Open Pit Dewatering Equipment

 

The dewatering activities will include the following:

 

horizontal drain holes in bench faces

 

sloped pit floors as required

 

in-pit sumps

 

vertical dewatering wells

 

a dewatering tunnel behind the north Mitchell Pit high wall

 

water collection system

 

Pit water will be collected and transported to the WSF.

 

Open Pit Support Equipment

 

The mine support equipment fleet requirements are summarized in Table 16.8.

 

The fleet size in Year 10 is shown as representative of the LOM requirement.

 

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Table 16.8Mine Support Equipment Fleet

 

Fleet Function Year 10
Track Dozer – 430 kW Shovel Support and Dump Maintenance 6
Rubber Tired Dozer – 350 kW Pit Clean Up 2
Fuel/Lube Truck Shovel and Drill Fueling and Lube 4
Wheel Loader Multipurpose – 14 t Pit Clean Up 4
Water Truck – 20,000 gal Haul Roads Water Truck 5
Motor Grader – 400 kW Road Grading 7
Tire Manipulator Tire Changes 4

 

Open Pit Ancillary Equipment

 

The mine ancillary equipment fleet includes such equipment as excavators, mine rescue trucks, picker trucks, cranes, snowcats, forklifts, service trucks, welding trucks, sump pumps and float trucks.

 

Snow Fleet

 

All of the following snow fleet equipment is chosen to start operating during pre-production and continue to the end of mine life, unless otherwise noted.

 

Six 37 t scrapers are included in the fleet. The scrapers are required to haul and spread crushed rock for traction control and remove snow from the haul roads and mine working areas as necessary. The scrapers are also used on occasion for small earthmoving jobs and reclamation projects outside of the snow season.

 

One wheel loader with an approximately 14-t bucket to clear snow from the plant area and truck shop, as well as ancillary routes within the mine. The wheel loader is also used to load the cone crusher at the crushing and screening plant.

 

Six snowcats to transport operators to equipment in a location that is inaccessible to the crew bus or vans because of heavy snowfall.

 

The snow fleet has a low utilization as it is only required in winter. Other than the use of the scrapers for summer construction projects and stockpiling road crush, operating the snow fleet equipment outside winter is not currently scheduled.

 

Open Pit Ancillary Facilities

 

Shops and Offices

 

In addition to providing an area for maintenance bays, tire shops, and a wash bay, the maintenance shop will also house:

 

a welding bay

 

an electrical shop

 

an ambulance

 

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a first aid room

 

a first aid office

 

a machine shop area

 

a mine dry

 

a warehouse

 

offices for administration, mine supervision, and engineering/geology staff

 

a lunchroom and foreman’s office.

 

The recommended shop sizing for the open pit operations includes eight service bays, one welding bay, and three wash bays. This will accommodate the fleet for the LOM PFS production plan. The mine maintenance facility will also include a machine shop area, tool storage area, mine muster area, warehouse, and office complex. A separate tire bay facility will be required with an exterior heated pad to accommodate at least two trucks and a tire manipulator.

 

16.3Mine Production Schedule

 

The summarized production schedule results are shown in Table 16.9 and Figure 1.2. The mine production plan starts in higher grade pits combined with lower strip ratios (higher margin pit phases) as much as possible to reduce pre-stripping costs and time. Development mining within the pit limits will start with the development fleet, transitioning to the large-scale equipment fleet. As work space becomes available. Pre-stripping will use the owner’s personnel and mining fleet in the pre-production years and will expose sufficient ore to ensure the ongoing production can sustain the mill requirements onward after mill startup. After mill startup, cut-off grade strategy is used to enhance revenues for a minimum capital payback period. The cut-off strategy stockpiles lower grade open pit material early in the mine life. The LOM strip ratio is 1.05.

 

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Table 16.9Summarized Production Schedule

 

    Year -2 -1 1 2 3 4 5 6 7 8 9 10 11-20 21-30 31-33 LOM
Pit to Mill Amount Mt     38 38 71 20 46 25 0 6 9 35 601 638 154 1,681
Au g/t     0.88 0.91 1.12 1.08 1.00 0.76 0.00 0.90 0.78 0.80 0.62 0.54 0.67 0.65
Cu %     0.23 0.23 0.20 0.21 0.19 0.30 0.00 0.05 0.05 0.24 0.12 0.12 0.15 0.14
Ag g/t     2.5 2.9 3.6 3.7 3.1 3.4 0.0 1.8 1.7 1.1 2.1 2.1 2.0 2.2
Mo ppm     44 40 49 53 51 38 0 144 123 126 95 73 72 79
Pit to Stockpile Amount Mt     46 66 142 57 19 11 9 24 34 29 75 109 0 622
Stockpile Reclaim Amount Mt     0 10 0 51 25 47 71 65 63 36 111 74 59 611
Au g/t     0.00 0.79 0.00 0.85 0.94 0.73 0.67 0.65 0.62 0.73 0.60 0.40 0.31 0.62
Cu %     0.00 0.18 0.00 0.21 0.25 0.19 0.16 0.19 0.20 0.30 0.09 0.12 0.06 0.16
Ag g/t     0.0 2.0 0.0 3.1 4.4 2.5 2.4 2.1 1.8 1.1 1.3 2.5 1.3 2.1
Mo ppm     0 84 0 47 27 62 69 76 76 92 71 71 69 69
Stockpile Balance Amount Mt     46 103 244 250 244 208 146 105 77 70 35 70 11 11
Total Mill Feed Amount Mt     38 47 71 71 71 71 71 71 71 71 712 712 213 2,292
Au g/t     0.88 0.89 1.12 0.92 0.98 0.74 0.67 0.67 0.64 0.77 0.62 0.52 0.57 0.64
Cu %     0.23 0.22 0.20 0.21 0.21 0.23 0.16 0.17 0.18 0.27 0.12 0.12 0.12 0.14
Ag g/t     2.5 2.7 3.6 3.3 3.6 2.8 2.4 2.0 1.8 1.1 2.0 2.1 1.8 2.2
Mo ppm     44 49 49 49 43 54 69 82 81 109 92 73 71 76
Metal to the Mill Au million oz     1.1 1.4 2.6 2.1 2.2 1.7 1.5 1.5 1.5 1.8 14.1 12.0 3.9 47.3
Cu million lb     197 233 317 334 335 357 257 273 286 424 1,844 1,888 576 7,320
Ag million oz     3 4 8 7 8 6 5 5 4 3 45 48 12 160
Mo million lb     4 5 8 8 7 8 11 13 13 17 144 115 33 385
Waste Mined (including rehandle) Amount Mt     216 211 87 12 82 78 45 30 19 25 787 820 6 2,417
Pre-production Waste Moved Amount Mt 247 287 0 0 0 0 0 0 0 0 0 0 0 0 0 534
Landbridge Reclaim Tonnes Amount Mt 0 0 0 0 0 0 0 0 0 0 0 0 0 187 243 430
Total Waste Moved Amount Mt 247 287 216 211 87 12 82 78 45 30 19 25 787 1,007 249 3,381
Total Material Mined Amount Mt 0 0 300 315 300 89 146 113 54 60 62 89 1,463 1,567 160 4,720
Total Material Moved Amount Mt 247 287 300 315 300 89 146 113 54 60 62 89 1,463 1,754 403 5,684

 

Note:

1Waste mined in the production schedule in Figure 1.2 includes re-handled waste and waste mined from borrow pit sources for construction purposes.

2The mill feed specified in Table 16.9 only includes ore from the Proven and Probable open pit Mineral Reserves and does not include any Inferred Mineral Resources.

 

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Figure 16.12 KSM Mill Feed Production Schedule

 

 

Source: MMTS (2022)

 

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17.0Recovery Methods

 

17.1Introduction

 

The proposed KSM plant for the 2022 PFS will start with nominal process rate of 130,000 tpd and expand to 195,000 t/d by Year 3 of operations. The process plant will receive ore from Mitchell, Sulphurets, and East Mitchell deposits. The planned mill life is approximately 33 years, excluding the production development stage and closure stage. The Mitchell deposit will be the dominant resource of mill feed for the process plant and will supply mill feed throughout the projected LOM. Ore from Sulphurets deposit will be fed to the plant together with the ores from Mitchell and East Mitchell pits starting from Year 5.

 

A combination of conventional flotation and cyanidation processes are proposed for the PFS. The processing flowsheet was developed based on the test results discussed in Section 13. In general, the mineralization from Mitchell, Sulphurets, and East Mitchell deposits are amenable to the combined flotation and cyanidation process. The process plant will consist of three separate facilities:

 

an ore primary crushing and handling facility at the mine site

 

an ore transportation system by trains through the MTT

 

a main process facility in the Treaty OPC area at the plant site, including following process facilities:

 

qcoarse ore stockpiles

 

qsecondary crushing by cone crushers in closed circuit with screens

 

qtertiary crushing by HPGRs in closed circuit with screens

 

qprimary grinding by ball mills

 

qcopper-gold/molybdenum bulk flotation

 

qregrinding of bulk concentrate, followed by bulk concentrate cleaner flotation

 

qcopper-gold/molybdenum separation depending on molybdenum grade of mill feed

 

qcopper-gold concentrate and molybdenum concentrate dewatering

 

qgold cyanide leaching of bulk copper-gold scavenger cleaner tails and copper removed pyrite rougher concentrate

 

qgold recovery by carbon elution and production of gold doré

 

qcyanide recovery, and cyanide destruction of washed leach residue prior to disposal of the residue in the lined pond within the TMF.

 

HPGR is proposed for the tertiary crushing because of its energy efficiency. The process is shown in the simplified flowsheet in Figure 17.1 and detailed in the following sections.

 

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Figure 17.1 Simplified Process Flowsheet

 

Source: Tetra Tech (2022)

 

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17.2Major Process Design Criteria

 

The concentrator is designed to process an average of 195,000 t/d. The major criteria used in the design are shown in Table 17.1.

 

Table 17.1Major Design Criteria

 

Criteria Unit Value
Average Daily Process Rate t/d 195,000
Operating Year d 365
Primary/Secondary Crushing
Availability – Primary Crushing % 70
Availability – Secondary Crushing % 85
Primary Crushing Product Particle Size, P80 mm 150
Secondary Crushing Product Particle Size, P80 mm <45
HPGR/Grind/Flotation/Leach
Availability % 94
Milling and Flotation Process Rate t/h 8,644
Mill Feed Size, P80 mm 2.0
Primary Grind Size, P80 µm 125-150
Bond Ball Mill Work Index - Design kWh/t 16
Bond Abrasion Index g 0.293
Concentrate Regrind Size, 80% Passing
Cu/Au Rougher/Scavenger Concentrate µm 15-20
Au-Pyrite Concentrate µm 20
Gold-bearing Materials Leach Method - Cyanidation
Feed Mass to CIL Circuit t/h 1,117

 

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Figure 17.2 Treaty Process Plant Layout

 

Source: Tetra Tech (2022)

 

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17.3Process Plant Description

 

The process plant is designed to process 195,000 t of ROM per day in three separate lines, excluding bulk cleaner flotation, copper molybdenum separation, and gold and silver extraction circuits which will be operated in one line. In the initial two years, the designed mill feed rate is 130,000 t/d. Additional equipment will be installed before Year 3 to increase the mill process rate to 195,000 t/d in Year 3. The processing plant layout at Treaty site is shown in Figure 17.2. The electricity power required for the process plant will be from FLT1 located at southeast of the process plant. Details of the power supply are described in Section 18.11.

 

17.3.1Primary Crushing

 

At the Mitchell OPC site, primary crushing will consist of three 60 in by 89 in gyratory crushers, three apron feeders, one train load surge bin feed conveyor, and associated transfer conveyors. The ROM material feeding the gyratory crushers will be mainly from the Mitchell pit in the initial years with main supplement from the upper zone of East Mitchell. Although the ore from the Sulphurets pit will be fed to the mill in Year 5, its main contribution will occur during Year 9 to Year 12. Most of the ore from the main East Mitchell zone will be introduced to the mill after Year 13. The ROM will be approximately 80% passing 1,200 mm. The oversized materials will be fragmented by rock breakers. The gyratory crushers will reduce the ROM material to a particle size of 80% passing 150 mm or less. The products from each gyratory crusher will be fed to one 1.83 m wide by 37 m long conveyor via one 2.13 m wide by 10 m long apron feeder. The crushed ore from the feeders will be transferred and fed to a 2.13 m wide by 495 m long train load surge bin feed conveyor, which will be located inside of the Mitchell surge bin feed conveyor tunnel. The surge bin is designed to have a live capacity of 30,000 t (two pockets, each 15,000 t).

 

The ROM ore from the Sulphurets pit will be trucked to the Mitchell site and crushed at the Mitchell crushing facility. Similarly, the ore from the East Mitchell deposit will be trucked to the three crushers for crushing to 80% passing approximately 150 mm.

 

17.3.2Coarse Ore Transport From Mitchell Site to Treaty Site

 

The crushed ore will be reclaimed by two automatic train loading systems from the two coarse ore surge bins located underground at the Mitchell site and transported to the plant site at Treaty site by a train transport system through the MTT. Loading chutes under the ore surge bins will feed ores into awaiting trains that will transport the ores to an unloading station at the Treaty end of the MTT. The train cars will dump ore into a live underground unloading bin. Two apron feeders will unload the bin onto two conveyors to transport the ore to the two coarse ore stockpiles adjacent to the MTT tunnel port at the Treaty site. Loading chutes will be controlled remotely and unloading chutes will operate autonomously. No onboard operators will be required within the tunnels during train system operation. Because the loading and unloading systems including surge bins are located underground, the arrangement would mitigate potential freezing and weather interruption issues.

 

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Each train will consist of two, 140-t electric locomotives and eighteen 42-m3 belly dump ore cars that will deliver the ore from Mitchell to Treaty. On average, 11 trains will deliver approximately 195,000 t/d of ore to meet the process plant requirements. An additional two trains will be on standby to provide for mechanical availability or to handle an increase in plant feed of up to 12,000 t/h. The transport system is detailed in Section 18.4. Dust collecting systems will be installed at the loading and unloading points to collect fugitive dust.

 

17.3.3Coarse Material Handling

 

The crushed ore from the trains will be continuously and automatically unloaded from the bottom discharge train cars into the bin underneath, with each car taking an average of nine seconds to unload. The train will be driven via traction drives across the unloading station at a maximum speed of 2.5 km/h.

 

At the bottom of the surge bin, the ore will be reclaimed by two apron feeders and then onto two conveyor belts that will transport the ore to the surface and feed two Treaty coarse ore stockpiles, each with a live capacity of 90,000 t at the Treaty OPC covered site. Apart from the conveyor tunnel, a vertical escape tunnel that joins the unloading station and the surface will be constructed for emergency egress.

 

The ore will be reclaimed from the stockpiles, each stockpile by six 1.4 m wide by 8.5 m long apron feeders and conveyed in two lines to the secondary crushing circuit. A total of four conveyors will reclaim the crushed ore to four double-deck vibrating screens for secondary crushing. A dust collecting system will be installed to collect fugitive dust generated at each of the transfer points. The reclaim tunnels will be heated to prevent potential freezing during operation in winter.

 

17.3.4Secondary Crushing

 

The reclaimed coarse ore will be conveyed to the secondary crushing facility and fed to four vibrating screens. Each screen oversize will feed a secondary cone crusher. The cone crushing facility consists of four cone crusher feed surge bins and four cone crushers. Each secondary crusher is in closed circuit with a screen. The cone crusher product will return to the screen feed conveyor. In normal operation, all four cone crushers will be in operation. If one cone is down for maintenance, the other three crushers will be capable of delivering the required tonnage for the downstream operation.

 

Screen undersize product that are finer than 50 mm will be delivered by conveyor to three HPGR feed surge bins, each with a 3,250-t live capacity feeding two HPGR crushers. The circuit will consist of the following key equipment:

 

four cone crushers, each with an approximately 3.4 m diameter mantle and driven by an 1,865-kW motor or equivalent,

 

four 3.7 m wide by 7.3 m long double deck vibrating screens (one additional unit on standby), and

 

related conveying system.

 

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17.3.5Tertiary Crushing Material Conveyance/Storage

 

The crushed ore from secondary crushing will be conveyed to three tertiary crusher HPGR feed surge bins, each with a live capacity of 3,250 t.

 

The ore from the HPGR surge bins will be further crushed by six HPGR crushers. Six belt feeders will withdraw the ore from the three HPGR feed surge bins and feed each of the six HPGR crushers separately. Each HPGR crusher is in a closed circuit with a 4.0 m wide by 8.0 m long double deck vibrating screen. Discharge from the HPGR crushers will be conveyed to three 3,250-t ball mill feed surge bins prior to being fed to the six ball mill feed screens at a controlled rate. The material is wet screened at a cut size of approximately 6 mm. The screen oversize will return to the HPGR feed bin while the screen undersize will leave the crushing circuit and report to the ball mill grinding circuits. The six HPGR crushing lines will have a total process capacity of 8,644 t/h. The key equipment is as follows:

 

six HPGR crushers, each equipped with two 2,900 kW motors

 

six 4.0 m wide by 8.0 m long vibrating screens

 

six 1.5 m wide by 10.0 m long belt HPGR feeders

 

six 1.4 m wide by 8.5 m long belt feeders for ball mill feed screens

 

related conveying system.

 

17.3.6Primary Grinding

 

The grinding circuit will employ conventional ball mills to grind the HPGR product to a particle size of 80% passing 125 to 150 µm. All the primary grinding circuits are designed to have a nominal processing rate of 8,644 t/h, each line with a nominal capacity of 2,881 t/h.

 

The primary grinding circuit will include six milling circuits (four circuits for the initial 130,000 t/d at Year 1 and six circuits for 195,000 t/d at Year 3), which are made up of the following equipment:

 

six 7.9 m diameter by 14.0 m long (26 ft. by 46 ft.) ball mills, each mill driven by two 7.5 MW synchronous motors

 

nine 600 mm by 500 mm centrifugal slurry pumps (six in operation and three on standby), each equipped with a 1,678 kW variable speed drive

 

six hydrocyclone clusters, each with twenty 710 mm diameter hydrocyclones.

 

Each ball mill will be in a closed circuit with a cluster of hydrocyclones. The hydrocyclone underflow will gravity-flow to the ball mill feed chute, while the overflow of each hydrocyclone cluster with a solid density of approximately 35% weight/weight (w/w) will gravity-flow to one of the six copper-gold-molybdenum bulk rougher flotation trains.

 

Lime will be added to each mill as required. Flotation collectors will be added to the hydrocyclone feed sumps or to the hydrocyclone overflow collecting sumps.

 

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17.3.7Copper, Gold and Molybdenum Flotation

 

Copper-Gold/Molybdenum Bulk Rougher/Scavenger Flotation

 

There will be three copper-gold-molybdenum bulk rougher flotation circuits. The overflow from two of the six hydrocyclone clusters from the primary grinding circuits will separately feed the three flotation trains, each consisting of five 500 m3 flotation cells. The flotation reagents used will include lime, A208, 3418A, fuel oil, carboxy methyl cellulose (CMC) and MIBC. A bulk copper-gold/molybdenum rougher flotation concentrate, approximately 6% to 7% of the flotation feed by weight, will be reground. The flotation tailing will be sent to the pyrite flotation circuit.

 

Copper-Gold/Molybdenum Bulk Concentrate Regrinding

 

The bulk concentrates from the three copper-gold/molybdenum rougher flotation circuits will be combined and reground to a particle size of 80% passing 15 to 20 µm in a two-stage regrind circuit. The first stage regrind circuit consists of three tower mills each with an installed power of 2,240 kW, in a closed circuit with a 250 mm diameter hydrocyclone cluster. The overflow of the hydrocyclones with a particle size of 80% passing approximately 30 to 35 µm from the first stage regrind mills will further be classified by a 150 mm diameter hydrocyclone cluster. The hydrocyclone underflow will be pumped to two stirred mills for further regrinding. The second regrind mill discharge, together with the second stage hydrocyclone overflow, will be pumped to the bulk copper-gold/molybdenum cleaner circuit.

 

Copper-Gold/Molybdenum Bulk Concentrate Cleaner Flotation

 

The hydrocyclone overflow will be cleaned in three stages. In the first stage of cleaner flotation, six 160 m3 tank cells will be used; for the second and third stages, three 100 m3 tank cells and one 100 m3 tank cell will be used separately. First cleaner flotation tailing will be further floated in two cleaner scavenger flotation cells, each with 160 m3 capacity. The concentrate product from the cleaner scavenger flotation will be sent to the first cleaner cells and the tailing will report to the gold leaching circuit. The tailing from the second and third cleaner flotation stages will be returned to the head of the preceding cleaner flotation circuit. Final copper-gold/molybdenum bulk concentrate will be sent to copper-gold/molybdenum bulk concentrate thickener.

 

The same reagents used in the rougher flotation circuit will be employed in the cleaner flotation circuits.

 

Copper-Gold and Molybdenum Separation

 

Depending on molybdenum content, the final copper-gold/molybdenum concentrate may be further processed to produce a copper-gold concentrate and a molybdenum concentrate. The separation will employ a conventional process, which will include copper suppression by sodium hydrosulphide and four stages of molybdenum cleaner flotation and regrinding. The circuit will include the following key equipment:

 

one 18 m diameter high-rate thickener

 

six 50 m3 conventional mechanical flotation cells

 

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one 1.5 m diameter by 4.5 m high column cell

 

one 1.1 m diameter by 4 m high column cell

 

two 1.0 m diameter by 4 m high column cells

 

one nitrogen gas generator

 

one regrinding stirred mill.

 

The copper-gold/molybdenum bulk concentrate will be thickened prior to the copper-gold/molybdenum separation. The thickener underflow will be diluted and conditioned with sodium hydrosulphide and gravity flow into the molybdenum rougher flotation cells. The rougher flotation tailing will be scavenged by flotation and the scavenger concentrate will return to the rougher flotation head while the tailing will be the final copper-gold concentrate reporting to the copper-gold concentrate dewatering circuit.

 

The resulting rougher molybdenum concentrate will be classified by a hydrocyclone. The hydrocyclone underflow will be reground by a stirred mill. The regrind mill discharge will join with the hydrocyclone overflow and report to the molybdenum cleaner flotation circuit. Four stages of cleaner flotation were designed to upgrade the molybdenum rougher flotation concentrate to marketable grade. The tailing of each cleaner flotation will be returned to the head of the preceding molybdenum cleaner flotation circuit while the first cleaner tailing will be sent to the molybdenum rougher flotation conditioning tank. To reduce sodium hydrosulphide consumption, the molybdenum flotation cells will be aerated by nitrogen gas, which will be generated on site by a nitrogen generator.

 

The final cleaner flotation concentrate will be leached to reduce the copper content if it is higher than 0.2%. The leached product will be dewatered in a molybdenum concentrate dewatering facility.

 

17.3.8Concentrate Dewatering

 

The upgraded copper-gold concentrate, together with copper sulphide precipitate generated from the cyanide recovery circuit by SART treatment, will be thickened in an 18 m diameter high-rate thickener. The thickener underflow will be directed to the copper-gold concentrate pressure filter to further reduce the water content to 9% moisture. The copper-gold concentrate will be stockpiled on site and then transported by trucks to a port facility at Stewart, where the concentrate will be stored and loaded into ships for ocean transport to overseas smelters.

 

The average annual output of copper concentrate for the initial two years, including ramp-up period in Year 1 and a lower nominal process rate of 130,000 t/d in Year 2, will be 345,000 dmt/a. After completion of process plant expansion to 195,000 t/d in Year 3, the annual copper concentrate production is expected to increase to approximately 505,000 dmt/a (Year 3 to Year 10 average).

 

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The molybdenum concentrate will be dewatered using a process similar to the copper-gold concentrate dewatering. The filtered concentrate will be further dewatered by a dryer to 5% moisture content, before being bagged and transported to the processors. The key equipment used in the dewatering processes will include:

 

copper-gold concentrate dewatering:

 

qone 18 m diameter high-rate thickener

 

qone 8 m diameter by 7 m high concentrate stock tank

 

qtwo 240 m2 pressure filters.

 

molybdenum concentrate dewatering:

 

qone 2 m diameter high-rate thickener

 

qone molybdenum concentrate leaching system

 

qone 4 m2 pressure filter

 

qone 2.5 t/h dryer.

 

17.3.9Gold Recovery From Gold-bearing Pyrite Products

 

Gold-bearing Pyrite Flotation

 

The tailing of the copper-gold/molybdenum rougher flotation circuits will be further floated in a pyrite flotation circuit. The pyrite rougher flotation will consist of three parallel lines, each line with five 500 m3 pyrite rougher flotation cells.

 

Tailing from the pyrite rougher flotation will gravity-flow or be pumped to the TMF located southeast of the main process plant.

 

Gold-bearing Pyrite Concentrate Regrinding

 

The pyrite concentrate will be reground to a particle size of 80% passing approximately 20 µm in four 2,240 kW tower mills. A hydrocyclone cluster consisting of thirty-six 250 mm diameter hydrocyclones will be incorporated with the mills in a closed circuit. The hydrocyclone overflow will report to the gold leach circuit or the copper-pyrite separation circuit to recover liberated residual copper minerals.

 

Depending on copper content, the reground materials may be subjected to a flotation process to separate copper minerals from the other minerals. The copper concentrate will be sent to the copper-gold/molybdenum cleaner flotation circuit, while the flotation tailing will report to the gold leach circuit.

 

Gold Leach

 

The reground gold-bearing pyrite product and the first cleaner scavenger tailing from the copper-gold/molybdenum bulk flotation circuit will be thickened to a solids density of approximately 65% in one 65 m diameter high-rate thickener.

 

The underflow of each thickener will be pumped to one direct cyanide leaching circuit, consisting of two aeration pre-treatment tanks and nine cyanide leaching tanks. In the pre-treatment tanks, the thickener underflow will be diluted with barren solution to approximately 45% w/w and aerated. Lime will be added to increase the slurry pH to approximately 11.

 

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The pre-treated slurry will be leached by sodium cyanide to recover gold and silver in a conventional direct cyanidation circuit. The leaching circuit will consist of nine agitated tanks, which are 14.5 m diameter by 14.8 m high. Six leaching tanks will be installed in Year 1 and three additional tanks will be installed in Year 2 for the increased throughput from 130,000 t/d to 195,000 t/d. Compressed air will be sparged under the agitators.

 

The leach residue will be dewatered in one 65 m diameter high-rate thickener. The thickener underflow with a solids density of 65% w/w will be diluted to approximately 40% w/w and further leached in a CIL circuit. The overflow will be treated by a SART process to recover copper and cyanide, and gold and silver adsorption to recover the dissolved gold and silver in a bank of four 3.5 m diameter by 5.5 m high carbon columns.

 

The CIL circuit consists of eight 14.5 m diameter by 14.8 m high CIL leaching tanks (three of the tanks will be installed later in Year 2). Similarly, compressed air will be sparged under the agitators.

 

The loaded carbon leaving the first tank of the CIL leaching bank will be transferred to the carbon stripping circuit, while the leach residue will be blended and sent to subsequent processes including residue washing, cyanide recovery, and cyanide destruction circuits.

 

The key equipment in the leach circuit will include:

 

two 65 m high-rate thickeners

 

two 14.5 m diameter by 14.8 m high aeration tanks

 

nine 14.5 m diameter by 14.8 m high direct leaching tanks

 

eight 14.5 m diameter by 14.8 m high CIL leaching tanks, each equipped with in-tank carbon transferring pumps and screens

 

one 2.5 m wide by 3.5 m long carbon safety screen.

 

Compressed air will be provided for the leaching process from four dedicated oil-free air compressors.

 

Cyanide and Copper Recovery

 

The overflow of the direct leach residue thickener will be sent to a cyanide recovery circuit where the copper and the cyanide will be recovered from the solution by a SART process. The post SART solution will be subjected to gypsum removal treatment prior to being sent to the gold adsorption circuit.

 

The SART treatment includes the leach solution acidification by sulphuric acid, copper, and other heavy metal precipitation from the solution by sodium hydrosulphide. The copper sulphide precipitates will be blended with the copper-gold concentrate for sale.

 

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The overflow of the CIL leach residue washing thickener will be sent to a separate cyanide recovery circuit where the copper will be removed, and the cyanide will be recovered from the solution by a SART/AVR process. The SART treatment is similar to the one used for the solution from the direct cyanidation circuit. The resulting copper sulphide precipitates will be blended with the copper-gold concentrate for sale. A part of the SART treated solution containing free cyanide released from copper and other heavy metal complexes will be directly recycled back to the CIL circuit while the balance is further treated by AVR cyanide recovery process. The solution will be pumped to two volatilization towers in series which will be carried out in a negative pressure system generated by a vacuum system. The solution will be sprayed from the top of the tower and pass through a fixed bed matrix while stripping air is blown from lower section of the tower to provide a high liquid surface area to promote volatilization.

 

The gas phase will be directed through an absorption tank, in which a caustic solution is circulated counter-current to the gas to absorb hydrogen cyanide. The regenerated cyanide solution will be returned to the leach circuit.

 

The cyanide-depleted solution from the volatilization tower will be circulated to the leach residue washing circuit and the leach circuit after the solution is treated with lime to a pH above 9.5.

 

Carbon Stripping and Reactivation

 

The loaded carbon will be treated by a Zadra pressure stripping process for gold desorption.

 

The loaded carbon will be transferred to two elution vessels. The stripping process will include the circulation of the barren solution through a heat recovery heat exchanger and a solution heater. The heated solution will then flow up through the bed of the loaded carbon and overflow near the top of the stripping vessels. The pregnant solution will flow through a back pressure control valve and then be cooled by exchanging heat with the barren solution prior to reporting to the pregnant solution holding tank for subsequent gold recovery by electrowinning. The barren solution from the electrowinning circuit will then return to the barren solution tank for recycling.

 

The stripping process will include barren and pregnant solution tanks, two 3.5 t stripping vessels, four heat exchangers, and two solution heaters and associated pumps.

 

Prior to reactivation, the stripped carbon will be screened and dewatered. The reactivation will be carried out in an electrically heated rotary kiln at a temperature of 700°C. The activated carbon will be circulated back to the CIL and CIC circuits after abrasion treatment and screen washing.

 

The carbon reactivation process will include a reactivation kiln, a carbon quench tank, and a carbon abrasion tank equipped with an attrition agitator, a reactivated carbon sizing screen, a carbon storage bin, and fine carbon handling associated equipment.

 

The reactivated carbon may be washed by acid in two acid washing columns to remove the metals/impurities that are loaded onto the carbon during gold and silver adsorption prior to being reused in the leaching circuits.

 

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Gold Electrowinning and Refining

 

The pregnant solution from the elution system will be pumped from the pregnant solution stock tank through electrowinning cells where the gold and silver will be deposited on stainless steel cathodes. The depleted solution will be subsequently reheated and returned to the stripping vessel. The electrowinning circuit will have a capacity to process 150 kg/d of gold-silver doré bullion and will include two 3.5 m3 electrowinning cells, direct current (DC) rectifiers, cathodes, anodes, and a pressure filter.

 

Periodically, the stainless steel cathodes will need to be cleaned to remove precious metal residues by pressure washing. The cell mud will fall into the bottom of the electrowinning cells and pumped through a pressure filter for dewatering on a batch basis. The filter cake will be transferred to the gold room for drying and smelting after it is mixed with melting flux. A 125-kW induction furnace will be used for gold-silver refining. The area will be monitored by a security surveillance system.

 

17.3.10Treatment of Leach Residue

 

Leach Residue Washing

 

The residue from the CIL circuit will be pumped to a two-stage conventional CCD washing circuit. The CCD circuit will consist of two 65 m diameter high-rate thickeners. The thickener overflow from the first stage washing will be pumped to the cyanide recovery system. The underflow (washed residue) of the second thickener will be sent to the cyanide destruction circuit prior to being pumped to the TMF.

 

Cyanide Destruction

 

The remaining cyanide in the washed leach residue from the second washing thickener will be decomposed by a sulphur dioxide (SO2)/air oxidation cyanide destruction process. Sodium metabisulphite will be used as the sulphur dioxide source. The equipment will include a 6 m diameter by 6 m high pre-aeration agitation tank, three 12.5 m diameter by 13.5 m high sulphur dioxide oxidation tanks, and a wet alkaline scrubbing system. Compressed air will be provided for the oxidation process. The treated residue will be sent to the copper removal treatment circuit.

 

Copper Removal

 

A copper removal circuit is proposed to further remove the dissolved copper from the treated residue slurry if the copper level in the slurry from the sulphur dioxide/air cyanide destruction circuit is higher than the requirement. Activated carbon will be added to the residue slurry after the slurry is treated by cyanide destruction. The copper removal treatment will be carried in two stages in two reactors. The loaded carbon will be removed from the first stage of the copper removal reactor, while the fresh carbon will be added into the section stage of the copper removal reactor. The copper loaded carbon will be stripped by acid washing and the copper in the washing solution will be precipitated by sodium sulphide. The precipitate produced will be blended with the copper-gold concentrate for sale. The treated residue will be sent to the lined CIL residue cell in the TMF.

 

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17.3.11Tailing Management

 

The flotation tailing and the treated CIL residue will separately gravity-flow to a tailing sand preparation facility adjacent to the TMF located southeast of the main process plant. Then both the tailing and CIL residue will continuously flow to the TMF or be pumped to the TMF depended on the dam crest elevations. The flotation tailing and CIL residue will be stored in separate areas within the TMF. If needed, the flotation tailing will be classified by hydrocyclones to separate coarse tailing fraction for dam construction which will mainly be in operation in summer.

 

The CIL residue will be deposited in a lined CIL Residue Cell. The residue will be covered with the supernatant to prevent oxidation of the sulphide minerals. The residue will be eventually covered by the flotation tailing. The supernatant from the CIL residue pond will be reclaimed by pumping to the CIL circuit for reuse. The excess water will be sent to a hydrogen peroxide (H2O2) water treatment plant to further remove impurities before it is sent to the north or south tailing cells.

 

There will be three flotation tailing pipelines directing the flotation tailing to the TMF. A standby pipeline will be installed for emergency bypass or maintenance requirements. If required, the flotation tailing will be classified to produce coarse tailing sands by two stages of hydrocyclone classification. The coarse fraction will be used to construct the tailing dam and the fines will directly report to the TMF. The supernatant from the tailing impoundment area will be reclaimed by two reclaim water barges and sent to the process water tanks by two stages of pumping. The reclaimed water will be used as process water for flotation circuits.

 

One energy recovery system will be installed on one of the rougher flotation tailing lines to generate electricity.

 

A separate barge equipped with reclaim water pumps will be installed in the flotation tailing storage pond to release the excess water via the Treaty Creek diffuser. Discharge will occur during a five-month window, beginning during spring runoff when the creek flows are highest. A floating skimmer will be installed. If required, flocculant will be added from the floating skimmer to improve the settlement of any suspended solids before the excess water is discharged.

 

17.3.12Reagents Handling

 

The reagents used in the process will include:

 

Flotation: PAX, 3418A, A208, PE26 (a mixture of CMC, fuel oil, MIBC, lime (CaO), NaHS, and sodium silicate (Na2SiO3)

 

Cyanidation and Gold Recovery: lime, sodium cyanide (NaCN), activated carbon, sodium hydroxide (NaOH), hydrochloric acid (HCl) and flux

 

Cyanide Recovery and Destruction Reagents: metabisulphite (MBS), copper sulphate (CuSO4), sulphuric acid (H2SO4), lime, NaOH, activated carbon, sodium hydrosulphide (NaHS)

 

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Others: flocculant, antiscalant, H2O2.

 

All the reagents will be prepared in a separate reagent preparation and storage facility in a containment area. The reagent storage tanks will be equipped with level indicators and instrumentation to ensure that spills do not occur during operation. Appropriate ventilation and fire and safety protection will be provided at the facility.

 

The liquid reagents (including fuel oil, A208, 3418A, PE26, MIBC, HCl, H2SO4, H2O2, and antiscalant) will be added in the undiluted form to various process circuits via individual metering pumps. PE26 will be added as needed only to control the effect of slimes on flotation.

 

All the solid type reagents (including PAX, NaHS, Na2SiO3 if required, NaOH, NaCN, CuSO4, and MBS) will be mixed with fresh water to 10% to 25% solution strength in the respective mixing tanks, and stored in separate holding tanks before being added to various addition points by metering pumps.

 

Lime will be slaked, diluted into 15% solid milk of lime, and then distributed to various addition points through a closed and pressurized loop.

 

Flocculant will be dissolved, diluted to less than 0.2% strength, and then added to various thickener feed wells by metering pumps.

 

Flux will be added directly in solid form.

 

17.3.13Water Supply

 

Three separate water supply systems will be provided to support the operation: a fresh water system, a process water system for grinding/flotation circuits, and a process water system for cyanidation and gold/silver recovery circuits.

 

On average, based on the preliminary water requirement and water balance estimates, approximately 59,000 m3/day of process make-up water and 2,100 m3/day of fresh water will be required for the processing plant.

 

Fresh Water Supply System

 

Fresh and potable water will be supplied to two 12 m diameter by 9 m high storage tanks from nearby wells and local drainage runoff areas. One tank will be located at the plant site and the other at mine site. Fresh water will be used primarily for the following:

 

fire water for emergency use

 

cooling water for mill motors and mill lubrication systems

 

potable water supply

 

reagent preparation.

 

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By design, the freshwater tanks will be full at all times and will provide at least 2 h of firewater in an emergency. The minimum freshwater requirement for process mill reagent preparation is estimated to be approximately 90 m3/h on average.

 

The potable water from the fresh water source will be treated (by chlorination and filtration) and stored in a covered tank prior to delivery to various service points.

 

Process Water Supply System

 

Two process water systems will supply the process water for the process plant. The water for each circuit will be from different sources, as follows:

 

Water for Grinding/Flotation Circuits: reclaimed water from the flotation tailing pond, copper-gold/molybdenum concentrate thickener overflow and the CIL feed thickener overflow, as well as fresh water. The dominant process water will be the supernatant fluid from the flotation tailing impoundment area.

 

Water for CIL Leaching/Gold Recovery Circuits: reclaimed water from the CIL storage pond, barren solution, and fresh water. As required, the water reclaimed from the flotation tailing pond may also be used in these circuits.

 

The water reclaimed from the flotation tailing impoundment area will be sent to two 25 m diameter by 15 m high process water surge tank by two stages of pumping systems, while the bulk concentrate thickener overflow will be directed to the primary grinding circuits. The process water tanks will be located approximately 25 m higher than the process plant base elevation. The water will flow to the various service points by gravity. A booster pump station is provided at the plant site to pump water to the various distribution points where high pressure water is required.

 

The water from the CIL Residue Cell will be pumped to an 8 m diameter by 8 m high process water surge tank located at the plant site. The water will service the CIL leach/gold recovery circuits. Any excess water from the CIL Residue Cell will be treated at the H2O2 WTP located at the Plant Site. The treated water will be sent to the north or south tailing ponds.

 

The overall site water management is detailed in Section 18.2.

 

17.3.14Air Supply

 

Plant air service systems will supply air to the following areas:

 

flotation circuits – low-pressure air for flotation cells by air blowers

 

leach circuits – high-pressure air by dedicated air compressors

 

cyanide recovery and destruction circuits – low-pressure air for cyanide recovery by dedicated air blowers and high-pressure air by dedicated air compressors

 

filtration circuit – high-pressure air for filter pressing and drying of concentrate by dedicated air compressors

 

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crushing circuit – high-pressure air for the dust suppression (fogging) system and other services by an air compressor

 

plant service air – high-pressure air for various services by two dedicated air compressors

 

instrumentation – instrument air at mine site and plant site will come from the local air compressors and will be dried and stored in dedicated air receivers.

 

17.3.15Assay and Metallurgical Laboratory

 

The assay laboratory will be equipped with necessary analytical instruments to provide routine assays for the mine, process, and environmental departments.

 

The metallurgical laboratory, with laboratory equipment and instruments, will undertake all necessary test work to monitor metallurgical performance and to improve the plant production and metallurgical results.

 

17.3.16Process Control and Instrumentation

 

The plant control system will consist of a Distributed Control System (DCS) with PC-based Operator Interface Stations (OIS) located in the following three control rooms:

 

Mitchell site primary crusher control room

 

MTT train transport control room

 

Treaty plant site control room.

 

The plant control rooms will be staffed by trained personnel 24 h/d.

 

An integrated remote operation system has been incorporated into the study. The integrated remote operations center (IROC) will be located offsite to remotely control and operate mining, process and water treatment plant operations. The IROC will be operated by trained personnel 24 h/d. The integrated operation is the integration of people, organizations, work processes, and information/operational technology to make faster and smarter decisions. It is enabled by global access to real-time information, collaborative technology, and integration of multiple expertise across disciplines, organizations, and geographical locations.

 

In addition to the plant control system, a closed-circuit television (CCTV) system will be installed at various locations throughout the plant including the crushing facility, the stockpile conveyor discharge point, the slurry pumping tunnel, the tailing facility, the concentrate handling building, and the gold recovery facilities. CCTVs in these areas will be monitored from the local control room and the central control room.

 

An automated train dispatching system will be utilized to achieve a safe and efficient flow of trains through the system, with no on-board operators. The system employing full radio-based train spacing and speed supervision on the whole railway system will be supervised from a control room located in the train maintenance shop. The train control system will operate using a wireless communications system (Wi-Fi) that must be in place for the entire track. While wireless communications are the current state of the art technology for train control communications, it is recognized that more efficient and reliable communications may be developed in the future.

 

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Process control will be enhanced with the installation of an automatic sampling system. The system will collect samples from various streams for on-line analysis and the daily metallurgical balance.

 

To protect operating staff, cyanide monitoring/alarm systems will be installed in the cyanide leaching area as well as at the cyanide recovery area and destruction area. A sulphur dioxide monitor/alarm system will monitor the cyanide destruction area as well.

 

17.4Yearly Production Projection

 

In general, the mineral processing is designed to use conventional flowsheet and mature technologies for the process plant. The flowsheet proposed is relatively simple and mining will start with conventional open pit operation with an exposed ore body on the surface. HPGRs will be used for comminution circuit, which is expected to have fewer potential rock hardness issues.

 

It is estimated that the plant may take approximately 12 months to reach 130,000 t/d after the plant is wet commissioned.

 

According to the metallurgical projections described in Section 13.6 and the current mine schedule, metal recovery and concentrate grades for the plant operation life are projected on a yearly basis, as indicated in Table 17.2. The flotation concentrate grades for the low copper head grade materials has been adjusted to a minimum concentrate of 23% Cu by adjusting copper, gold, and silver recoveries to achieve the target grade.

 

As shown by the test results, it is anticipated that, on average, the impurity contents in the copper concentrates would be below the penalty limits as outlined for most of the smelters, although in short periods the impurity content may slightly exceed the penalty limits as outlined for some of the smelters. The projected copper concentrate quality is shown in Table 17.3.

 

In general, the molybdenum concentrate separated from copper and molybdenum bulk concentrate will be leached on site to remove copper, lead, and other impurities. The anticipated molybdenum content is approximately 50%. The main impurities such as copper and lead are estimated to be lower than 0.2% and 0.3%, respectively.

 

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Table 17.2Projected Metallurgical Performance

 

 

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Table 17.3Projected Copper Concentrate Quality

 

Element Unit Content
Range Years 1 to 7
Average
LOM
Average
Cu % 23 – 26 25 24
Au g/t 38 – 96 74 65
Ag g/t 38 – 318 210 180
Mo % 0.03 – 0.21 0.15 0.10
ST % 28 – 41 34 34
S-2 % 26 - 36 32 32
Fe % 24 – 35 29 29
Sb ppm 100 – 2,200 850 800
As ppm 200 – 4,500 1,800 1,400
Bi ppm 10 – 170 47 40
Hg ppm 0.5 – 10 2 2
F ppm 65 - 480 150 150
Cl % 0.01 – 0.02 0.01 0.01
Se ppm 50 – 140 50 85
Pb % 0.1 – 1.0 0.2 0.3
Zn % 0.2 – 1.8 0.4 0.5
SiO2 % 2.3 – 11 6.0 6.0
CaO % 0.2 – 1.0 0.5 0.5
Al2O3 % 0.5 – 4.0 2.0 2.0
MgO % 0.1 – 0.5 0.3 0.3
MnO % 0.01 – 0.04 0.02 0.02
InSol % 3.1 – 12 - -

 

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18.0Project Infrastructure

 

18.1Site Layout

 

There will be two separate areas of infrastructure associated with the 2022 PFS: the Mine Site and the PTMA. The Mine Site is the centre of mining activity and includes the primary crushing facilities (Mitchell OPC). Process facilities will be located at the Treaty OPC in the PTMA, approximately 23 km northeast of the Mine Site. Twinned tunnels (the MTT) will be constructed from multiple excavation headings to connect the mine site to the PTMA. Along the MTT route, a topographical low (valley) is designated as the Saddle Area, approximately 17 km from the Mitchell portal, where the MTT will be accessed via a construction adit. The TMF is located in a valley comprising the upper catchments of North Treaty and South Teigen creeks, southeast of, and adjacent to the Treaty OPC.

 

The Mine Site layout at the end of construction is shown in Figure 18.1.

 

The Treaty OPC area is shown in Figure 17.2 and the TMF area is shown in Figure 18.2.

 

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  Figure 18.1Mine Site Layout after Initial Construction

 

Source: Tetra Tech, 2022

 

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  Figure 18.2Ultimate PTMA Layout

 

Source: Tetra Tech, 2022

 

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18.2Tailings, Mine Rock, and Water Management

 

18.2.1Introduction

 

This section addresses the geotechnical designs for tailings and mine rock management, as well as for site-wide water management. Designs were advanced in 2022 to incorporate incremental improvements and changes to water management structures in the Mine Site and TMF areas, in response to commitments made during the EA process review.

 

2016 design improvements in the TMF area (KCB, 2016b) included:

 

addition of a discharge pipeline system and a diffuser located in Treaty Creek to route operational-period discharges to Treaty Creek

 

rerouting of the North Cell Closure Spillway, and

 

relocation of the TMF seepage collection dams downstream to more effectively intercept seepage and to align with permit conditions.

 

2022 design improvements in the TMF area included:

 

primary cyclones will be located in a single house near the North Dam

 

secondary cyclones will be located on their respective dam crests (North Dam, Splitter Dam, and Saddle Dam), and

 

conceptual mitigations to address potential cycloned sand supply shortfalls.

 

2022 PFS design improvements at the Mine Site included:

 

relocating the MDT outlet to an area of reduced geohazards and to support improved year-round construction access. Revised tunnel slope to eliminate need for exposed surface spillway

 

updating the MDT and MTDT alignments and designs as single axis, partially-lined or fully-concrete-lined tunnels

 

including an inlet gate incorporated into the MDT alignment to control inflows during maintenance and repair

 

adjusting the North Pit Wall Dewatering Adit (NPWDA) alignment and timing to reflect updated LOM Mitchell Pit shells

 

staging of open pit production, allows surface diversion of Mitchell valley until Year 3, when the Mitchell Valley Drainage Tunnel (MVDT) would be commissioned

 

aligning the McTagg non-contact closure channel with an updated Mitchell RSF geometry.

 

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In 2016, KCB re-assessed additional site climate and hydrology data recorded through 2015. These analyses determined similar values to those adopted for the 2012 PFS (KCB, 2012a; KCB, 2013a; Tetra Tech, 2012) for an average year. As a result, the water management design basis remained unchanged from 2012.

 

18.2.2Mine Site Characterization

 

Other than 2019 Mitchell Valley rockfill sourcing investigations for the WSD, 2021 hydrogeological drilling in the WSD, and 2021 foundation condition testing for temporary water treatment in Mitchell Valley, no additional geotechnical site investigations have been completed for the WSD or RSF areas since the 2012 PFS (Tetra Tech, 2012).

 

In 2019 geomechanical and geophysical testing was completed for monzonite rockfill for the WSD from the P8 Quarry which has similar properties to the Sulphurets Pit (KCB, 2019b).

 

In 2021 WSP Golder completed a hydrogeological drilling program in the WSD area (Golder, 2021).

 

Foundation condition testing for Mitchell Valley temporary water treatment facilities was completed by KCB (KCB, 2021).

 

The 2012 PFS provides a results summary of previous KCB Mine Site geotechnical site investigations completed up to 2012 (KCB, 2009; 2010; 2011; 2012b; and 2012c).

 

Seabridge has completed regional geological mapping that resulted in a Mine Site geology map, presented in Section 7.

 

Mine Site Climate

 

Much of the annual precipitation at the Mine Site and PTMA occurs as snowfall between October and May, while peak rainfall is associated with storms coming in from the Pacific between August and October. Major elevation variations and numerous glaciers help create diverse climatic conditions across the site.

 

Two climatic regions are present within the site development areas: the western region in the Sulphurets watershed (Mine Site) and the eastern region in the Treaty-Teigen watersheds (PTMA). The two regions are 23 km apart and have differing climates. The two areas are separated by the Johnstone Icefield (ranging from 1,800 m to 2,200 m in elevation).

 

Significant orographic and rain shadow effects were recorded in the KSM area as part of the 2012 baseline study. In 2012, KCB and ERM performed extensive analysis of climate variations in Mine Site and PTMA for engineering design and EA purposes (Rescan, 2013). Algorithms were developed based on the UBC watershed model to estimate effects of variation in precipitation with altitude, and to adjust glacier and snow melt rates in response to climatic variations.

 

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In 2016, metrological and hydrological data collected since 2012 was reviewed by KCB (KCB, 2016d). The result is that no significant trends have been observed in the additional data for average site parameters, as compared to the data prior to 2012.

 

Mine Site Temperature

 

Weather data recorded at the Sulphurets weather station between 2007 and 2015 indicate the following:

 

mean annual temperature is approximately 0°C

 

mean monthly temperatures range from -13°C in December to 14°C in July

 

temperature extremes range from -31°C to 30°C

 

mean daily temperatures are above freezing from May to October and below freezing from October to May.

 

Mine Site Precipitation and Hydrology

 

The estimated mean annual precipitation is 1,652 mm at Sulphurets weather station (elevation of 880 masl). Annual lake evaporation is estimated at 400 mm. Runoff at the Mine Site is influenced by the effects of both seasonal snowmelt and glacial melt. Mitchell and McTagg glaciers are losing significant ice mass on an annual basis. Runoff from glacier-influenced catchments is therefore larger than the annual precipitation over these catchments. Effects of glacial meltwaters are included in the analysis of flows and extreme events. Monthly precipitation, evaporation, and runoff distribution for the Mine Site, as well as the runoff distributions for the glacier catchments, is provided in Table 18.1.

 

Precipitation listed in Table 18.1 is representative of the Sulphurets weather station at 880 masl and is derived from site data between 2008 and 2011. Additional site data recorded since 2012 has not changed the values adopted for design.

 

The 2012 and 2016 KCB design reports present detailed analyses of climate and hydrology data for the Mine Site (KCB, 2012b; KCB, 2016b) and PTMA (KCB, 2012a; KCB 2016b).

 

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  Table 18.1Climate Data for the Mine Site (Sulphurets Creek Climate Station)

 

Data Area Jan Feb Mar Apr May Jun Jul Aug Sep Oct Nov Dec Total
Precipitation (mm) 1 215 50 50 66 99 83 115 149 264 297 132 132 1,652
Pond Evaporation (mm)2 0 0 0 0 86 93 99 80 43 0 0 0 400
Site Runoff Distribution (%) 0 0 1 1 4 14 35 17 17 7 3 1 100
Mitchell Glacier Runoff (mm) 37 37 73 73 110 404 1,248 917 514 183 37 37 3,670
McTagg Glacier Runoff (mm) 66 33 33 33 197 625 954 526 461 197 99 66 3,290
Notes:1Weather station at 880 masl elevation.
 2Estimated pond evaporation based on pan evaporation data.

Source: KCB

 

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Mitchell Glacier Recession

 

Seabridge has been monitoring the recession of site glaciers by analyzing historical air photos, Light Detection and Ranging (LiDAR) data, ongoing GPS and remote sensing measurements of the glacier extents. Glaciers in the Mine Site area continue to recede; recession rates for the Mitchell Glacier toe area recession since 2008 have reached as much as 65 m/a and total recession has exceeded 2012 estimates. As a result, the Mitchell pit area is now ice free. The locations of the initial stage proposed surface contact water inlets in the toe area of Mitchell Glacier are now also free of ice.

 

18.2.3TMF and PTMA Site Characterization

 

TMF Site Investigations

 

No additional geotechnical site investigations have been conducted at the TMF since the effective date of the 2016 PFS, with the exceptions of a hydrogeological drill program by WSP Golder and a program of mapping and sampling for till borrow material (KCB, 2019a) and a 2021 program that included refraction seismic and geotechnical/hydrogeological drilling and testing in the PTMA plant site area (KCB, 2021)

 

A hydrogeological site investigation in the TMF area was conducted by WSP Golder from June to September 2021, these investigations sought to reduce the hydrogeological data gaps at the TMF. Twelve boreholes were conducted in the TMF area. Instrument installations included thirteen vibrating wire piezometers, at six locations, while ten monitoring wells were installed at ten locations (Golder, 2021).

 

Tailings Characterization and Laboratory Testing

 

No additional tailings testing has been conducted for TMF design purposes since 2012. The Treaty process plant will produce two tailings streams: the bulk rougher flotation tailings1 representing about 90% of the ore (by dry weight) and a fine, sulphide-rich cleaner tailings comprising 10% of the ore. The sulphide stream will be cyanide leached using the CIL method and then processed for gold recovery. A two-stage cyanide destruction circuit is proposed, using the Inco sulphur dioxide process, followed by hydrogen peroxide treatment2.

 

The flotation tailings is classified as NPAG and will be cycloned to produce sand fill for construction of the tailings dams during the summer months. The fine cyclone overflow tailings will be discharged along the upstream crest of the tailings dams. The entire flotation tailings stream will be discharged along the dam crests during the winter months.

 

The CIL residue tailings is a high-sulphide concentration material and is classified as PAG. This material will be deposited under water in the CIL Residue Storage Cell in the centre of the TMF and kept saturated to mitigate the onset of acid generation.

 

 

1Referredto as “Flotation Tailings” in this report.
2Thisstream is referred to as “CIL Tailings” in this report.

 

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TMF Area Climate

 

Additional climate and hydrological data collected since 2012 was reviewed (KCB, 2016d) and the conclusion is that no significant trends in average parameters are apparent. As a result, the climate and hydrology parameters used for design of the TMF in 2012 have been retained.

 

TMF Area Temperature

 

Weather data recorded at the TMF area from the Teigen Creek weather station between 2009 and 2011, with additional data now available through 2015, shows the following:

 

mean annual temperature is approximately 0°C

 

mean monthly temperatures range from -8°C from December to February, to 11°C in July

 

temperature extremes range from -27°C to 29°C

 

mean daily temperatures are above freezing from May to October and below freezing from October to May.

 

The monthly precipitation and runoff distribution are provided in Table 18.2.

 

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  Table 18.2Climate Data for the TMF (Teigen Creek Climate Station)1

 

Data Area Jan Feb Mar Apr May Jun Jul Aug Sep Oct Nov Dec Total
Precipitation (mm) 151 110 123 82 55 69 82 82 165 206 164 82 1,371
Pond Evaporation (mm)2 0 0 0 0 75 81 86 70 38 0 0 0 350
Site Runoff Distribution (%) 1 1 1 3 16 32 19 8 9 7 2 1 100
Note:1Weather station at elevation 1,085 masl.
 2Estimated pond evaporation based on pan evaporation data.

Source: KCB

 

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18.2.4Rock Storage Facilities

 

A rendering of the ultimate mine site layout, including the Mitchell RSF is provided in Figure 18.3. Design of the RSFs has not changed since the 2016 design other than the elimination of the McTagg RSF and inclusion of a closure channel on the Mitchell RSF to replace the McTagg Closure Channel.

 

At the PFS level, there are three primary RSF design considerations:

 

foundation conditions

 

maximum lift height

 

closure slope criteria.

 

Conservative RSF designs were developed in collaboration with MMTS to address the aforementioned design considerations using existing data. MMTS designed the RSF layouts, with geotechnical guidance on slope stability and geotechnical recommendations from KCB.

 

The RSFs will be built in progressive lifts (bottom-up construction) to initially confine toe areas and consolidate foundations to improve stability and reduce downslope risks.

 

18.2.5Mine Site Water Management

 

Figure 18.3 illustrates ultimate water management structures as existing at the end of mine life showing diversion tunnel routes and operational phase surface diversions.

 

Due to elimination of Mitchell Block Cave the MDT design criteria is reduced from a 1,000-year to 200-year return period. This reduction is justified by the lower sensitivity of an open pit to flood flows compared to a block cave. As a result, the MDT alignment can be a single axis. Additionally, the MDT is re-aligned to avoid mineralized areas and to reduce tunnel length. The normally wetted depth of the revised MDT is concrete-lined to improve reliability, reduce erosion and to reduce potential for loadings from areas of mineralized wall rock. The revised MDT alignment improves reliability and constructability of the diversion with reduced risk from geohazards. The revised tunnel slope and alignment precludes MDT hydropower development due to the challenging terrain for powerhouse and power transmission line development.

 

Elimination of the Mitchell Block Cave also allows for shortening and realignment of the NPWDA that carries contact water from east of Mitchell Pit. Glacier recession and the elimination of the inlet setback, that was previously required to avoid block cave subsidence cracking, allowed replacing the two additional stages of the surface inlets shown in prior studies with a single stage.

 

McTagg Valley catchments are diverted around Mitchell RSF and the WSF by an inlet at the same location as in the approved EA. The MTDT alignment discharges into the EA outlet location in Gingrass Creek. The MTDT’s design is a single bore design that is fully concrete-lined. This design reduces tunneling volume, improves reliability, and matches the rock types along the route. The MTDT includes provision for a single-stage, fully-developed hydroelectric facility from start-up, instead of only in the later stages as in prior studies, thus replacing the start-up hydroelectric capability that was eliminated at the MDT. The redesigned MTDT hydroelectric facility has a larger catchment than in prior studies. The revised tunnel route and outlet portal location improves winter constructability and has direct access to the muck pads via a short construction access tunnel. The MTDT construction access tunnel also shortens the length of the penstock to the McTagg Power Plant.

 

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  Figure 18.3Mine Site Ultimate Water Management Facilities

 

Source: Tetra Tech, 2022

 

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After mining is complete, perimeter closure channels will be constructed at the top and margins of the RSF, and the channels widened for closure along the toes of the RSF. Operational phase channels are shown on Figure 18.3; closure routings are similar.

 

Diversion Tunnels and Surface Diversions

 

Locations of the MDT sub-glacial inlets have not been adjusted since 2012 (Tetra Tech, 2012). Elevations of the sub-glacial inlets, their arrangement, and the MDT alignment were adjusted for 2022 based on ice-penetrating radar and hot water ice drilling completed over Mitchell Glacier in 2019. The 2022 PFS MDT outlet has been shifted to a location that will support year-round construction via a construction access decline tunnel (KCB, 2022a).

 

The MDT alignment and elevation have been adjusted to avoid or perpendicularly intersect the main identified faults in the area based on information from exploration drilling. From the inlets, the tunnels curve eastward to intercept the Brucejack Fault at a near perpendicular angle. Further downstream, the alignment dips under the sub-horizontal Mitchell Thrust Fault. The MDT configuration, as of 2022 design updates, consists of a single, partially-lined tunnel.

 

Due to the weaker Stuhini Formation rock in the area, the MTDT alignment for the 2022 PFS will be a single, fully-lined tunnel. The inlet and outlet locations remain unchanged from the 2016 design.

 

Contact Water Collection Systems

 

The MVDT is a 5 km long, 5 m by 6 m tunnel that drains to the WSF and conveys contact water around the Mitchel RSF. Prior to the construction of the MVDT in Year 2, water will be routed around the pit and RSF area through surface channels to the WSF.

 

The MVDT connects to the NPWDA, which will be constructed and available by Year 5, to accept pit wall drainage and local drainage of contact water from upstream of Mitchell pit and from the Snowfields area.

 

The Mitchell RSF includes a Selenium Seepage Collection System that is designed to collect up to 500 L/s of seepage and convey this flow to the Selenium WTP. Water from other sources will be treated when seepage collected from the RSF is less than 500 L/s. The collection and treatment of seepage from the RSF, and other high selenium loading waters, will enable selective removal of selenium from flows with higher selenium concentrations, compared to the lower concentrations within the WSF. The Selenium Seepage Collection System will be operational by Year 5.

 

The Selenium WTP will discharge to the WSF to allow further treatment for metals removal.

 

Water Storage Facility

 

Seepage from the Mitchell RSF requiring treatment for the removal of metals by the HDS process will be collected in the lower Mine Site by the WSD. The WSD is an asphalt core rockfill dam that will create the WSF pond, and will be large enough to handle seasonal freshet flows as well as volume accumulated from a 200-year wet year.

 

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The estimated ultimate HDS water treatment capacity of 4.0 m3/s is greater than the annual average or monthly peak flows to allow the treatment rate to vary seasonally with stream flow rates.

 

The WSD design was updated after the 2012 PFS (Tetra Tech, 2012) that included dam slope and internal zonation revisions (KCB, 2012e). The slopes and zonation of the dam are shown in section view in Figure 18.4.

 

The WSD will be located in the lower Mitchell Creek area and founded on competent rock. The WSD crest elevation is also unchanged from the ultimate dam height in the current design and will be established at the full height of 716 masl (165 m height) before Year 1. An emergency spillway will be cut into rock on the southeast side of the dam. There will be appropriate freeboard for avalanche wave mitigation and flood routing.

 

Water in the WSF is predicted to be acidic, similar to existing seeps situated in the upper Mine Site. An asphalt core will be included in the dam to control seepage. Asphalt is inert with respect to acidic water. To control seepage, the WSD and WSF Seepage Dam foundations will be grouted. The depth of the WSD grout curtain is designed to vary from 25 m at the west abutment to as deep as 150 m at the east abutment if required. Grout hole spacing will be 2.5 m.

 

Fill for dam zones is specified such that critical zones of the dam (e.g., sections in contact with the core) and drain zones will be constructed with materials that have low potential to react with acidic water sourced from quarries tested to be geomechanically and geochemically adequate for construction. The anticipated quarry to be mined for WSD construction rock will be the P8 Quarry residing in the western area of the Mitchell pit on the north side of Mitchell Creek.

 

During the Application/EIS (Rescan, 2013) review process, additional mitigations were developed to minimize seepage from the WSD. These design enhancements are included in the 2016 PFS. Changes include six seepage interception tunnels that lower groundwater levels between the WSD and the WSF Seepage Dam to reduce the driving force on seepage. The seepage collection tunnels will also facilitate foundation grouting both during construction and for remedial grouting after completion of the WSD, if required.

 

An asphalt-core seepage collection dam will be located downstream of the WSF. Water collected in this dam will be sent to the WTP via an HDPE pipeline.

 

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  Figure 18.4Water Storage Dam Sections

 

Source: KCB

 

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The WSD Construction Diversion Tunnel (CDT) routes Mitchell Creek around the dam footprint during construction. The diversion tunnel size is designed as 4.4 m by 5.0 m and is sized to pass flows from a 50-year storm event. To reduce exposure to avalanche hazards, the downstream portal was moved upstream from where shown in the EA resulting in a total tunnel length of 900 m.

 

Hydroelectric Potential of Diversions

 

Diverting McTagg Creek into a tunnel creates an opportunity for hydroelectric power generation. Generated power can be used during mine operations or sold to the grid via the power lines through the MTT. During operations, the hydroelectric plant will reduce the power requirements of the mine. Upon mine closure, the hydroelectric plant will continue operating and will generate income and partially offset water treatment costs.

 

18.2.6Water Treatment

 

Temporary Mine Area Construction Period Water Treatment

 

Temporary Water Treatment Plants (TWTP) and settling ponds are designed to meet current guidelines for mine sediment control and settling ponds.

 

During the construction period, six TWTPs for TSS and metal removal will be provided in the proposed mine area. Additional TWTPs will be located in the Saddle Area, and at the Treaty portal of the MTT. The TWTPs are intended to deal with drainage from existing mineralized zones, PAG cuts, tunnel portals, and runoff from PAG tunnel muck piles during the period before the permanent WTP is in operation.

 

The WSF and WTP is intended to be in operation during the 18-month pre-production period to capture sediment and runoff from mine area stripping and from fill placement during Mitchell OPC and haul road construction.

 

The eight TWTPs and associated collection ponds will operate during the construction period to manage potential metals, TSS, and ammonia in drainage from tunnel portals and from temporary stockpiles of tunnel muck near the portals and other flows of contact water (Table 18.3).

 

  Table 18.3Temporary Water Treatment Plant Locations

 

TWTP Location
TWTP #1 WSD/HDS WTP Area
TWTP #2 MTDT Outlet
TWTP #3 MDT Outlet
TWTP #4 Saddle MTT Portal Tunnel
TWTP #5 MDT Inlet
TWTP #6 MTT – Mitchell Portals
TWTP #7 WSD CDT Inlet
TWTP #8 Treaty MTT Portals

Source: ERM

 

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High density Sludge Water Treatment Plant

 

The HDS WTP is designed to treat water that comes in contact with areas of disturbance from mining operations and natural seeps in the area. Design of the HDS WTP process is unchanged since the EA design. The HDS WTP was scaled down in 2022 due to revised estimated inflow rate commensurate with a reduction in impacted mine area.

 

Water will be collected in the WSF. Drainage from the Mitchell pit and Mitchell RSF will be directed by gravity to the WSF. The water from the WSF will be pumped over the WSD to the HDS WTP. The HDS WTP is designed with variable discharge rates in order to stage discharge to match the natural hydrograph, to ensure sufficient dilution capacity to minimize any effects on the receiving environment.

 

The HDS WTP maximum throughput capacity will be 4.0 m3/s; however, the maximum rate is only anticipated to be required during both freshet and autumn rains when receiving environment has its maximum naturally occurring flow. Water pumped from the WSF will pass through hydroelectric generators installed at the Energy Recovery Power Plant, immediately upstream of the HDS WTP.

 

The HDS WTP installed generation capacity will be 9 MW and the two installed turbines will be capable of passing a flow of up to 4.0 m3/s.

 

The site selection for the HDS WTP is based on a mine life and post-closure treatment for 200 years. The HDS WTP will be located at an elevation of 520 m on a flat benched terrain above the flood plain near the confluence of Mitchell and Sulphurets creeks.

 

During operation, the sludge will be transported year-round by truck to the Mitchell OPC and onto the ore trains within the MTT, where it will be transported with the ore to the stockpile located at the Treaty OPC, fed through the Treaty process plant, and deposited with the tailings in the TMF.

 

The three principal reagents for the HDS WTP will be quick lime, dry flocculant, and sulphuric acid for pH adjustment to 7.5. Table 18.4 provides an estimate of annual reagent consumption based on an annual average of 49.8 Mm3 of water treated. The predicted total annual volume of water will vary from 44 Mm3 to 53 Mm3. After closure, the predicted long-term volume of water for treatment will be 43 Mm3/a.

 

  Table 18.4Average Annual Reagent Consumption for the HDS WTP

 

Reagent Reagent Dosage Rate Average Annual
Water Treated
(Mm3/a)
Average Annual
Reagent Consumption
Quick lime to pH 10.5 0.49 kg/m3 49.8 25,000 t
Flocculant 1.79 g/m3 49.8 89 t
H2SO4 to pH 7.5 6.55 mL/m3 of 36.8N H2SO4 49.8 326 m3

Source: Tetra Tech

 

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Selenium Treatment Plant

 

Design of the Selenium Water Treatment Plant (Selenium WTP) is unchanged since the 2016 PFS. BioteQ Environmental Technologies Inc. (BioteQ) demonstrated selenium removal of spiked Mitchell Creek feed water during a pilot-scale ion exchange water treatment study (BioteQ, 2015).

 

The pilot study demonstrated reduction of selenium concentrations from 120 ppb and 320 ppb feed water to less than 1 ppb (BioteQ, 2015). The Selen-IX™ Ion Exchange Circuit is designed to selectively remove selenate from the feed water with a high efficiency in order to obtain the 1 ppb discharge limit, while concentrating the selenium into a small volume of brine solution that is directed to the eluate treatment circuit. The eluate treatment circuit removes selenium from the spent regenerant (or eluate) solution produced by the ion exchange circuit with an electro-reduction process using iron and fixes the selenium into an iron-selenium solid that is easily separated from solution. The solution discharged from the eluate treatment circuit is largely free of selenium and can be recycled back to the ion exchange circuit for re-use in resin regeneration.

 

As the Selenium WTP is only designed to remove selenium, effluent from the Selenium WTP will report to the WSF for further treatment at the HDS WTP, prior to discharge to the receiving environment.

 

18.2.7Tailings Management Facility Design

 

The general layout of the TMF is shown in Figure 18.2. The TMF will be located within a cross-valley between Teigen and Treaty creeks. Three cells will be constructed: the North Cell and the South Cell will store flotation tailings, and the CIL Residue Cell (fully lined with a geomembrane) will contain PAG CIL residue tailings.

 

The cyclone sand dams will be constructed over earth fill starter dams using the centerline construction method with compacted cyclone sand shells and low-permeability glacial till cores. The Saddle and Splitter dam cores incorporate geomembranes to limit seepage from the CIL residue tailings. The dams will be progressively raised over their operating life to an ultimate elevation of 1,068 m.

 

Seepage from the impoundment will be controlled with low-permeability zones in the tailings dams and dam foundation treatment. Seepage and runoff from the tailings dams will be collected downstream at seepage collection dams and pumped back to the TMF. The ponds behind the collection dams will also be used to settle solids eroded by runoff from the dam and fines from cyclone sand construction drain-down water.

 

Tailings Staging Plan

 

Tailings flows will be initially routed by gravity in slurry pipelines from the plant to the North Cell and CIL Residue Cell. Energy will be recovered during early years of operation of each cell from discharge of the tailings into the impoundment. Tailings will be pumped to the CIL Residue and South Cell when required during later stages of the operation.

 

Tailings will be retained by four cyclone sand tailings dams: the North Dam, Splitter Dam, Saddle Dam, and Southeast Dam. During operation, elevations of annual dam crest raises will be set to provide 12 months of tailings storage and to store the PMF plus 1 m of freeboard.

 

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Figure 18.5 illustrates the staging of the TMF. The North Cell will be constructed first and will store flotation tailings production for the first half of the LOM production plan. During operation of the North Cell, floods will be routed south. A pipeline and surface channel will divert environmental maintenance flows of up to 2 m3/s from the East Catchment around the TMF into Teigen Creek. After the North Cell has been filled and closed it will be reclaimed over a five-year period. The CIL Residue Cell will be constructed and operated in parallel with the North Cell, and in the first stage, will be filled to about half its capacity with PAG CIL residue tailings. Halfway through the LOM production plan, the South Cell goes into operation, providing flotation tailings storage for the remaining mine life. During this time, the East Catchment Tunnel will route East Catchment flood flows away from the South Cell. The CIL Residue Cell will be filled to ultimate capacity. The South Cell and CIL Residue Cell will then be closed and reclaimed over a five-year period.

 

Based on the mill ramp-up schedule, and assumed in-situ density ranges possible at start up (starting with an initial 1.0 t/m3 ramping up to 1.3 t/m3 by end Year 2), the starter dams can store about 18 months of tailings. The earth fill starter dams at the North, Splitter and Saddle dam sites will be constructed to store a minimum of 8.4 Mm3 of water for mill start-up. The design operating PMF ranges from 42 Mm3 at start-up to 91 Mm3 at the ultimate dam elevation. The dams will then be raised annually by cycloning tailings sand. The beach will be built up to separate the reclaim pond from the dams by at least 700 m, increasing to 1,200 m at the ultimate dam elevation. The separation between the tailings dam and pond created by the beach increases the margin of safety against overtopping of the tailings dam, and reduces seepage through the tailings dam and underlying foundation.

 

The South Starter Dam, at elevation 930 m, will be completed half way through the LOM production plan to allow deposition to begin in the South Cell.

 

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  Figure 18.5TMF Staging Plan

 

Note: Raising of cyclone dams within each stage is not shown on these diagrams.

Source: KCB

 

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TMF Dam Structures

 

Over an initial four-year construction period, three earth fill starter dams will be constructed at the North Cell and CIL Residue Cell (North, Splitter and Saddle) to provide start-up flotation and CIL residue tailings storage. Halfway through mine life the Southeast starter dam will be required to form the South Cell. These dams will be progressively raised over their operating life to an ultimate elevation, providing storage capacity of 2.29 Bt. A summary of the tailings dams is provided in Table 18.5.

 

TMF Starter Dams

 

Starter dams will be earth fill embankments, with shells of compacted random fill supporting the central glacial till cores. The glacial till cores will be keyed into the underlying foundations to cut off seepage through weathered near-surface soils and any pervious strata. A blanket drain is provided to lower the phreatic levels in the downstream shell. Riprap erosion protection will not be placed on the upstream slope due to the temporary exposure of the dam to the pond water. Figure 18.6, Figure 18.7 and Figure 18.8 show typical sections of the starter dams.

 

Main Tailings Dams

 

The North, Splitter, Saddle, and Southeast dams will be compacted cyclone sand dams with glacial till cores constructed by the centerline method. Dimensions of the dams are summarized in Table 18.5. Details of the North, Splitter, Saddle, and Southeast dam designs are shown in Figure 18.6, Figure 18.7, and Figure 18.8.

 

A system of finger drains will be installed at the base of the downstream shells of the North, Saddle and Southeast dams to keep water levels in the dam depressed. Main drains in the centre of the valley floor will collect and convey seepage to the toe of the dam. Smaller secondary drains will convey water laterally into the main drains.

 

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  Table 18.5Tailings Dam Summary

 

Dam Starter Dam Ultimate Dam
Crest
Elevation
(masl)
Maximum
Height*
(m)
Crest
Length
(m)
Random Fill
Volume
(Mm3)
Core
Volume
(Mm3)
Crest
Elevation
(masl)
Maximum
Height*
(m)
Ultimate
Crest
Length
(m)
Cyclone Sand
Volume
(Mm3)
Core Volume
above
Starter
(Mm3)
North Dam 930 80 680 3.59 0.95 1,068 218 1,900 47.16 3.42
Splitter Dam 935 61 890 3.74 1.08 1,068 194 1,930 31.31 3.75
Saddle Dam 935 35 780 2.09 0.75 1,068 168 1,600 22.99 3.39
Southeast Dam 930 101 890 12.32 1.72 1,068 239 1,400 60.45 3.20
Totals - - 3,240 21.73 4.50 - - 6,830 161.91 13.77

Note: *maximum height measured at dam crestline

Source: KCB

 

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  Figure 18.6North Tailings Dam

 

Source: KCB

 

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  Figure 18.7Saddle and Splitter Tailings Dams

 

Source: KCB

 

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  Figure 18.8Southeast Tailings Dam

 

Source: KCB

 

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TMF Dam Construction

 

Table 18.5 summarizes fill requirements for the dams. For construction of the starter dams, general fill and core material will be excavated by a contractor fleet from local borrow sources that were identified at each dam site. The cyclone sand TMF dams will be raised using a fleet of dedicated mine equipment.

 

TMF Seepage Recovery Dams

 

Seepage recovery dams will be constructed of compacted glacial till in a similar manner as for the tailings starter dams, but with flatter 3H:1V upstream and downstream slopes.

 

TMF Area Water Management

 

Figure 18.9 shows the schematic water balance with water inputs and outputs from the impoundment.

 

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  Figure 18.9Schematic TMF Water Cycle

 

Source: KCB

 

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TMF Diversion Channels

 

Two main diversion channels, the Northeast Diversion and the South Diversion, will be constructed around the TMF North Cell with additional diversions around the Treaty OPC to divert non-contact runoff water into a tributary of Teigen Creek at the north end of the TMF.

 

Diversion channels are shown on Figure 18.2.

 

TMF Area Extreme Flood Routing and Storage

 

The TMF cells are designed to be able to store extreme flood events without discharge. Specifically, the PMF can be stored, which is a flood resulting from a 30-day Probable Maximum Precipitation (PMP) storm, combined with a 100-year 30-day snow melt. The perimeter diversions are assumed to be inoperative during this extreme flood event.

 

During operations, water will be reclaimed from the ponds and routed back to the Treaty OPC, where it will be treated as part of the mineral separation process. Surplus water from the TMF will be discharged seasonally via the Treaty Creek Diffuser. Discharge will occur during an approximate period extending from May to mid-November, when the creek flows are highest.

 

18.3Tunnels

 

A number of tunnels will be excavated during both the pre-production period and the operating periods. These tunnels are classified as either infrastructure tunnels or water tunnels. Table 18.6 summarizes tunnels constructed in pre-production and Table 18.7 summarizes tunnels constructed in the operational phase.

 

The infrastructure tunnels provide for the transportation of ore, personnel, and supplies between the Mitchell Valley mining area and the Treaty OPC. The principal infrastructure tunnel is the MTT, which transports all mined ore from the Mitchell OPC to the Treaty OPC, and personnel and freight between the PTMA and the Mine Site via the train system. The MTT also includes electric power cables to service the Mitchell mining area.

 

The water tunnels include the diversion tunnels in Mitchell Valley for surface water management and the pit slope drainage tunnels for the Mitchell north high wall. At PTMA, in the later stages of TMF operations (after Year 17) a diversion tunnel is required to divert East Catchment flows once the South Cell is in operation.

 

This section includes a description of the construction method, sequencing, and cost basis for the tunnels as applied to the PFS schedule and capital estimate. The MTT design cross-section has been modified since the EA to accommodate the change to train haulage from the previous ore conveyor system and for the 2022 PFS the cross-section for the tunnels has been increased to accommodate a revised development fleet using 60 t low-profile haul trucks, rather than the previously planned rail based development haulage. This will speed up the development rate. The alignment remains essentially the same as previous studies with minor modifications to the portal locations and arrangements.

 

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The quantities below in Table 18.6 and Table 18.7 are based on revised tunnel lengths and cross sectional areas by SRK, KCB, and MMTS. For MTT related excavations, SRK provided information for the twin haulage tunnels (SRK, 2022) and MMTS added quantities for the other excavations included in the MTT system based on previous studies. For the water tunnels, quantities are based on KCB’s updated designs. (KCB, 2022b)

 

  Table 18.6KSM Pre-Production Tunnels Summary

 

Tunnel Description Total
Length
Excavation
Volume
Lining
Volume
    (m) (m3) (m3)
Infrastructure Tunnels - MTT
Principal Alignment Twin Tunnels 43,737 1,338,352  
  Saddle Declines 708 21,665  
  Ancillaries 15,933 487,550  
Treaty Ore Handling Conveyor Discharge Tunnel 514 30,000  
  Bulk Excavation   12,650  
  Cross-Tracks (Cross-Overs)  1,161 35,527  
  Mitchell Ore Handling  704 16,960  
  Bulk Excavation   22,057  
Infra. Tunnels Total:    62,757 1,964,761 0
Water Tunnels – MDT, MTDT, CDT and SCT
MDT Diversion Tunnel Inlets & other tunnels  2,295 59,934  
  Construction Access Tunnels Decline  1,830 45,750  
  Main Diversion Tunnels  6,570 145,854  
MDT Subtotal:   10,695 251,538 0
MTDT  Access Tunnel 210 3,780 32,370
  Main Tunnel 3,730 134,280 33,190
MTDT Subtotal:   3,940 138,060 65,560
Other Water Tunnels:        
CDT at WSD  Construction Diversion Tunnel  900 18,360  
SCT at WSD  Seepage Collection Tunnels  1,780 19,402  
Water Tunnels Total:   17,315 427,360 65,560
 
Pre-production Tunnels Total: 80,072 2,392,121 65,560
Note:Tunnel volumes are based on construction volumes and account for drill hole “look out” contractor tolerance where indicated includes concrete lining.
Source:Based on SRK (MTT haulage tunnels), MMTS (other MTT excavations) & KCB (water tunnels) and re-configured for costing.

 

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  Table 18.7KSM Operational Phase Tunnels Summary

 

Tunnel  Total
Length 
Excavation
Volume 
Lining
Volume
  (m)  (m3  (m3)
NPWDA 3,770 89,855 5,100
MVDT and MVDT/NPWDA Decline  4,000 131,240 2,560
East Catchment Tunnel at TMF 4,000 67,733 0
Operation Phase Tunnels Total  11,770 288,828 7,660
Note:Tunnel volumes are based on construction volumes and account for drill hole “look out” contractor tolerance where indicated includes concrete lining.
Source:Based on KCB (water tunnels) and re-configured for costing.

 

18.3.1Mitchell-Treaty Tunnels

 

The MTT is required to accommodate the ore transport system, which is an electric rail-based twinned haulage system to handle all ore from the Mitchell mining areas to the Treaty OPC (see Figure 18.10) It also handles personnel transport, supplies, and services between the two operating areas. Personnel transport includes mine side shift exchange as well as daily personnel interchange. Supplies and services include rail-based fuel tanks, explosives, mine operating supplies, maintenance parts and supplies, and electric power cables. Any oversized cargo during operations will be brought to the mine area via CCAR.

 

The Mitchell-Teigen Tunnel Reserve 1006527 crosses Tudor Gold Corp.’s Treaty project mineral claims immediately northeast of KSM. Tudor Gold owns a 60% interest in the claims. In March 2021, Tudor Gold published the first resource estimate for the project. Seabridge and Tudor Gold are actively discussing a cooperative agreement towards mutually beneficial development of both properties.

 

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  Figure 18.10MTT General Location and Alignment

 

Source: SRK TER, 2022

 

MTT Design

 

Primary crushing of ore from the Mitchell OPC will produce material that will be transported through the MTT to the crushed ore stockpile located at the Treaty OPC, approximately 22 km to the east.

 

The twinned tunnel configuration provides higher capacity with the haulage loop and rail cross-overs allow sections of the tunnel to be isolated for periodic tunnel maintenance. Under normal operations, the North Tunnel will be designated for westbound travel and the South Tunnel will be designated for eastbound travel.

 

The MTT will comprise the following excavations:

 

twin tunnels, each approximately 21,900 m in length with seventy-three 30 m long cross-cuts, linking the two tunnels. Half of the cross-cuts will be equipped with refuge stations

 

nine 210 m long track cross-overs to allow for flexibility and tunnel maintenance during operations

 

two Saddle declines of length 305 m and 367 m, driven at a -15% gradient, will be used for ventilation and materials handling, respectively.

 

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other excavations include electrical substations to be located in the South Tunnel, sumps, remuck bays, truck loading bays, cover drilling bays and cut-outs for electrical starters.

 

MTT Tunnel Support and Advance Rates

 

The length of the tunnel has been broken into six general ground support classes to account for the different rock types and structures to be encountered in the tunnels, based on field investigations comprising mapping, drilling, and testing. These support class segments are accounted for the tunnel advance rates, and costs per meter for costing and scheduling.

 

MTT Mining Method

 

The tunnels will be constructed in accordance with BC Mines Act and Health, Safety and Reclamation Code for Mines in BC using conventional drill and blast techniques and will follow the conditions contained within the License of Occupation for the MTT issued in September 2014 by the Government of BC.

 

The MTT will be excavated by a contractor using conventional drill and blast with haulage to the portals via 60 t low-profile diesel-operating trucks. Face jumbos and rock bolting jumbos will be electric-hydraulic, therefore reducing ventilation requirements. Excavated rock will be temporarily stored at the portals in managed facilities and relocated to permanent rock disposal sites when the permanent waste management plan is in operation.

 

At the end of the MTT excavation, contractors will lay the track and install the electrical supply and distribution system for the trains.

 

Water Management

 

Both MTT tunnels will be driven with a ditch in the lower corner for the collection of tunnel water during the operations period. The nominal grade of the MTT is 1.46% down from the Treaty portals to the Mitchell portals; therefore, in operation the water in the tunnels will flow to the Mitchell portal where it will be routed to the WSF.

 

During construction, tunnel water will be collected and treated by TWTPs located near the portals.

 

MTT Mining Sequence

 

The MTT is on the construction critical path and has therefore been broken into two segments to allow for concurrent development workplaces resulting in a shorter total tunnel construction period. This preliminary sequence will be accomplished using the Saddle Declines, which are located in the Saddle area (See Figure 18.10). This is a transverse valley along the tunnel alignment, located approximately 5.3 km from the Treaty Portal of the MTT and 16.6 km from the Mitchell portal. Two declines will be driven at a negative grade from the Saddle and will access each of the north and south MTT tunnels. They will allow the Mitchell-Saddle segment of the MTT to be driven from two headings from either end of the segment, with four independent crews at one time. Additionally, two crews will be advancing the Treaty-Saddle segments from the Treaty portals, one in the North Tunnel and one in the South Tunnel.

 

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Crews from the shorter Saddle-Treaty segment of the MTT will finish early and will then be used to develop some of the other excavations that do not interfere with the advance of the primary excavations.

 

The North and South tunnels will advance together both from the Mitchell and Saddle headings. As the twin headings advance, cross-cuts will be developed every 300 m joining the two tunnels. The cross-cut closest to the face will be used for remucking, the next closest cross-cut will be used for the ventilation cross-over and the cross-cut before that will be equipped with a refuge station.

 

Ventilation During Construction

 

Since there are three sets of two headings advancing at one time, there will be three primary ventilation circuits. The primary circuit will be established by installing two fans in a bulkhead just inside the portal in the South Tunnel (and near the entrance to the Saddle adits), to provide fresh air under positive pressure through the South Tunnel, and through the ventilation cross-cut with exhaust out the North Tunnel.

 

The secondary circuits will be established to intercept fresh air from the primary circuits in order to ventilate the advancing faces. This will be done by two auxiliary fans with flexible vent ducting installed in the South Tunnel on the fresh air side of the active ventilation cross-cut and blowing air to the advancing headings in each of the South and North tunnels. The air from the South Tunnel will exhaust via the ventilation cross-cut where it will meet with the exhaust air from the North Tunnel. This exhaust air stream will then flow out the portal. As the tunnel faces advance, a new remuck cross-cut will be established and the previous remuck cross-cut will now act as the new ventilation cross-cut. The previous ventilation cross-cut will be sealed and equipped with the advancing refuge station.

 

Ventilation During Operations

 

During normal operations air will be moved through the tunnel by the piston effect created by the movement of the trains. Thus, ventilation in the North Tunnel will generally be from east to west and ventilation in the South Tunnel will generally be from west to east.

 

To allow for segments of the MTT to be isolated for maintenance, sets of ventilation doors with axial vane fans will be installed at the portals and at the track crossovers. In this way, fresh air will be supplied to the isolated sections of track where the crews will be working. Energizing or de-energizing of fans will be coordinated with train traffic so that the fans will not move air against each other or against closed vent doors.

 

18.4Mine To Mill Ore Transport System

 

At the Mitchell OPC, ore will be crushed and conveyed through a tunnel to two live underground bins within the MTT. Loading chutes under the bins will feed into trains that will transport the ore to an unloading station at the Treaty end of the MTT. The train cars will dump into a live underground unloading bin. Apron feeders will unload the bin onto a conveyor to transport the ore to the top of the Treaty COS.

 

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The ore transport is configured to start operations at a nominal production rate of 130,000 tpd and ramps up to 195,000 tpd for the start of Year 3. Each ore train Consist can carry approximately 1,600 tonnes per trip. Based on 20 hours per day and 350 days per year specifications for the two production rates are:

 

Ore Production Trains in operation at 130,000 t/d

 

Six trains, each consists of two 140 t Schalke Locomotives c/w 18 NMT 42m3 Bottom Discharge Wagons per consist.

 

Average speed 40 km/h

 

Cycle time 81.6 min

 

Productivity 1,167 t/h

 

Ore Production Trains in operation at 195,000 t/d

 

Nine trains, each consists of two 140 t Schalke Locomotives c/w 18 NMT 42m3 Bottom Discharge Wagons per consist.

 

Average speed 39 km/h

 

Cycle time 83.3 min

 

Productivity 1,144 t/h

 

Note: Two additional Consists have been included in Year 1 to account for reduced availability and as spare capacity to catch up production when required.

 

Speeds vary slightly for the different production levels due to traffic requirements from modelling.

 

Unloaded trains will be able to travel downwards with a speed up to 50 km/h. All traction motors are then operated in regenerative mode, feeding back approx. 870 kW (1.26% gradient) and 1580 kW (2.04% gradient) of power into the grid.

 

The transport system capacity has been confirmed at PFS level by simulation of the system using desktop simulation to define the train requirements.

 

The static modelling of the rail system has been carried out based on all trains travelling to the most remote part of the rail transport system. In practice, the average distance travelled by the trains will be substantially less than the distance to the furthest chute. Accordingly, the capacity of the system has been underestimated, providing some contingency for congestion and unplanned delays. There is room for refinement in future studies.

 

The trains will travel on a conventional ballasted track structure with treated timber ties and operate via an electrical overhead conductor rail system. Trains will be controlled by an automated train control system managed from a remote control room. Loading chutes will also be controlled remotely and unloading chutes will operate autonomously. Future developments in technology may enable full automation. No onboard operators will be required within the tunnels during train system operation.

 

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Figure 18.11 shows a schematic plan view of the approximately 21.9 km long MTT dual track transport system. The South Tunnel will be primarily utilized for loaded trains travelling to Treaty OPC, and the North Tunnel will be utilized for empty trains travelling back to Mitchell OPC. The tunnels will run uphill from Mitchell OPC to Treaty OPC at an overall 1.46 % incline, such that tunnel drainage will flow to the Mine Site. Cross-overs are planned at both end points of the tunnel, as well as in two intermediate points to split the route into three sections. One section of the tunnel can be isolated when maintenance is required, and one-way traffic can be implemented and safely controlled by the train automation system. A simulation verified that these traffic flow restrictions will not compromise average daily train production requirements. The chainage to the cross-overs will be determined during construction when in-situ conditions from tunneling activities and probe drilling indicate suitable locations for crossover intersections with the main tunnels without compromising the required traffic flow from the simulation.

 

Figure 18.11MTT Dual Track Plan View

 

Source: MMTS, 2022

 

Figure 18.12 shows a typical cross-section used for both of the MTT tunnels. The inner profile indicates the minimum dimensions for the operating trains. The outer profile indicates the larger dimension to facilitate the specified tunnel development equipment. The tunnel costs and productivities are based on the larger profile.

 

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Figure 18.12MMT Cross Section (Typical)

 

Source: SRK, 2022

 

Due to the 2.04% incline from Saddle to Treaty, two locomotives are required per train since a tractive effort of approx. 550kN necessary. The two locomotives combined require 4,800 kW of power at the wheels to maintain speeds between 45 km/h (1.26% gradient) and 38 km/h (2.04% gradient) while going uphill.

 

The current consumption of each locomotive is depending on the voltage level of the conduction rail system and the operating speed of the locomotives. Operating the loaded trains uphill with a maximum speed of 38 km/h on the 2.04% gradient will require a current draw of approximately 3,500 A for both locomotives on the train from the conduction rail at an operational voltage level of 1,500 V DC.

 

The locomotives, as well as the loading and unloading chutes, will carry their own fire suppression systems, and will not require a fixed system within the tunnel.

 

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An automated train dispatching system will be utilized to achieve a safe and efficient flow of trains through the tunnels. This will be located at the IROC to coordinate ore flow from the mine to the mill.

 

As shown in Figure 18.13, two parallel underground spur lines coming off the main tunnel will be used for loading ore at the Mitchell end of the MTT.

 

Trains will be continuously loaded at the Mitchell OPC without the need to stop between cars. Each 42 m3 car can be loaded in an average of 36 seconds. The mine ore cars adopted for this PFS are bottom discharge with an overlapping apron between the cars.

 

Mine ore cars will unload into the surge bin centrally under the unloading station. From the bottom of the bin, ore will discharge into two apron feeders and onto a conveyor belt that will transport the ore to the surface and feed to the Treaty COS.

 

18.4.1MTT Freight and Personnel Transport

 

The train system will also be used for transportation of personnel and freight between the Treaty and Mitchell areas via the MTT. Freight and personnel transport will be scheduled on a daily basis, and the transport trains will be controlled by the automated train control system.

 

Specially configured personnel and freight trains will transport personnel, freight, and fuel through the MTT, with marshalling and unloading areas at each end, separate from the ongoing ore transportation facilities. Personnel, freight, and fuel handling will only be scheduled during the day shift operations.

 

The train control system will ensure there is no haulage of fuel or explosives when personnel are being transported in the tunnels. In the event of an emergency, the personnel cars will be equipped with personal protective equipment (PPE) kits including self-rescuers. The twin tunnel configuration and cross-cuts provide alternate egress and rescue stations will be installed through the length of the MTT.

 

Treaty Staging Areas

 

Staging areas on surface near the Treaty Portal will be used to load personnel, freight, and fuel onto the specialty train cars. These staging areas will be road accessible and include a laydown area for all freight that is for transport through the MTT.

 

The personnel marshalling will be out of the South Tunnel and will include a structure to protect waiting passengers from the elements.

 

A train maintenance shop will be located in the freight marshalling area that will include workshop facilities, fuel storage and a control room for train system operations (see Figure 18.13).

 

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Figure 18.13Treaty Personnel, Freight, and Fuel Staging and Marshalling

 

Source: Tetra Tech, 2022 (Train Facility Design by NMT)

 

Mitchell Staging Areas

 

Three separate, enclosed, underground staging areas near the Mitchell portal will be used to offload passengers, freight, and fuel, respectively (shown in Figure 18.14). Loaded fuel tanks and other hazardous freight will be shuttled out of the tunnel as soon as possible. Personnel will exit the Mitchell portal by bus or other light vehicles.

 

Freight and fuel staging areas will include gantry cranes to offload the train payloads onto awaiting flatbed tractor-trailer units. Freight will be driven out to its ultimate destination at the Mine Site.

 

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Figure 18.14Mitchell Personnel, Freight, and Fuel Staging and Marshalling

 

Source: MMTS

 

18.5Site Roads

 

Currently, the KSM Site can only be accessed by helicopter. Helicopter support will augment the road pioneering work and construction camps set up. Avalanche mitigation will be constructed where appropriate so that work can be safely carried out in Mitchell Valley.

 

The Mine Site roads after initial construction and pre-production are shown in Figure 18.1. Figure 18.3 shows the Mine Site roads at the end of LOM.

 

18.5.1Road Width

 

Site road widths were designed to comply with the following BC Mines Regulations:

 

for dual lane traffic, a travel width of not less than three times the width of the widest haul vehicle used on the road

 

for single lane traffic, a travel width of not less than two times the width of the widest haul vehicle used on the road

 

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a berm height of at least three-quarters the height of the largest tire on any vehicle hauling along the road, where a drop-off of greater than 3 m exists.

 

Ditches are included within the travel width allowance.

 

18.6Ancillary Buildings

 

Ancillary building construction considered for the PFS will be pre-engineered, stick-built, or modular structures, as applicable. The HVAC for these buildings will be designed to industrial standards. The following ancillary buildings and infrastructure are included in the PFS:

 

Treaty OPC:

 

qfuel storage facility
qfuel distribution station
qadministration building
qassay and metallurgical laboratories
qwarehouse and maintenance building
qconcentrate storage/load-out building
qcold storage/reagent storage building
qfirst-aid building
qore train storage yard, maintenance shop and loading/unloading facilities
qpermanent operations camp
qpotable water treatment plant
qsewage treatment plant
qincinerator
qlandfill
qsubstation and auxiliary power supply facilities
qconstruction laydown areas
qpre-construction fuel storage
EPCM and contractors’ offices, concrete batch plant, temporary construction camps, TWTPs, and numerous other construction related facilities

 

Mitchell OPC and lower Mine Site areas:

 

qtruck shop including first aid facilities
qHDS WTP with sludge storage facilities
qSelenium WTP (operational by Year 5)
qdiesel fuel storage and dispensing
qpermanent operations camp
qtemporary construction camps
qsewage treatment plants
qincinerator

 

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qlandfill
off-site facility:

 

qnew concentrate storage and loadout at the Stewart port facility.

 

18.6.1Treaty OPC

 

Fuel Storage and Distribution (Permanent and Construction)

 

The main fuel storage tanks at the Treaty OPC are sized to store 1,300,000 L, which accounts for 6 days of fuel demand. All fuel storage areas will be lined with containment berms and approved double-wall type tanks. Additional fuel stations will be located near the Mitchell OPC, truck shop, and at the Sulphurets pit. Gasoline will also be similarly stored where required.

 

The majority of the fuel requirement is for mining activities at Mitchell; however, some fuel will be distributed to all Treaty OPC facilities via pipelines from the Treaty Fuel Storage Tank to the required facilities.

 

A pipeline from the Treaty fuel storage tank to the train marshalling area will be installed, with hook-ups for fast fuel transfer directly to the ISO fuel tanks at the marshalling area. Wherever possible, the ISO fuel tanks will not be removed from the train cars on the Treaty end.

 

Administration Building

 

The pre-engineered administration building will be approximately 1,000 m2 in plan area.

 

Assay, Metallurgical, Acid-base Accounting, and Geotechnical Laboratory

 

The pre-engineered laboratory will be an 815 m2 single-storey structure located in a separate building near the process plant at the Treaty OPC.

 

First Aid Buildings

 

The first aid buildings will be pre-engineered structures, located at both the Mine Site and Treaty OPC, equipped with first aid facilities and provide emergency vehicle storage.

 

Concentrate Storage

 

The on-site concentrate storage facility will be a pre-engineered structure, approximately 2,000 m2 in area. It will have a five-day storage capacity equating to approximately 4,600 t of concentrate.

 

Cold Storage/Reagent Storage Building

 

The cold storage/reagent storage building will be located at the Treaty OPC and will be approximately 1,200 m2 in area.

 

Permanent Camp

 

The Treaty operations camp will be located at the PTMA, approximately 600 m southwest from the process plant. The camp components will include accommodation, office/recreation complex, kitchen/diner, parking, sanitary sewer, potable water treatment and wastewater treatment.

 

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Warehouse and Maintenance Building

 

An 800 m2 warehouse and maintenance pre-engineered building will be constructed at the Treaty OPC. It will be located adjacent to the cold storage facility.

 

18.6.2Mine Site

 

Truck Shop

 

The Mine Site truck shop will be a pre-engineered building, approximately 9,500 m2 in area. The truck shop/mine dry will comprise eight maintenance bays, two light vehicle repair bays, a truck lube bay, a truck wash bay, a welding and machine shop, an electrical and instrument shop, a storage warehouse and a dry area.

 

Permanent Camp

 

The Mitchell operations camp will be located between Sulphurets and Gingras creeks, just north of the CCAR. Like the Treaty operating camp, the Mitchell operating camp is planned to be used for most of the construction period and the entire LOM.

 

The camp components will include a helipad, sleeping dorms, parking, fuel storage and loading area, recreation facility, sewage treatment, fire/fresh water tanks, an emergency generator, laundry, and kitchen/diner.

 

Landfills

 

Two landfills are proposed to be permitted and developed, one for the Mine Site and one for the PTMA.

 

The Mine Site landfill, which will occupy approximately 6.5 ha, will be located within the Sulphurets laydown area.

 

The PTMA landfill, occupying approximately 8.4 ha, will be located near the Treaty operating camp.

 

Each landfill will include a land farm. The land farms will accept contaminated soils from spill clean-ups and leaks, while the landfill will be used to dispose of non-inert, dry industrial, and forestry waste.

 

18.7Sewage

 

The wastewater treatment system installed at the construction and operation camps in both Mine Site and PTMA will treat the anticipated maximum daily flow through a variety of processes to meet a secondary level of treatment as defined in the Municipal Wastewater Regulation (MWR) (BC Reg. 87/2012) of the Environmental Management Act (2003).

 

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For temporary construction camps off site with occupancy less than 100, Type 3 effluent quality as defined in the Public Health Act (2008) sewerage guidelines will be applicable to the sewage requirements for these camps.

 

18.8Communications System

 

For permanent communications a fibre optic cable will be installed over the 30 km transmission line from the Treaty OPC to BC Hydro’s TCT substation beside Highway 37 near Treaty Creek. This fibre cable is required by BC Hydro for line protection and additional fibres have been allocated for communications. Tahltan Communications (a partnership between Tahltan Nation Development Corporation (TNDC) and CityWest Cable and Telephone Corporation) have a pair of fibres on the BC Hydro NTL transmission line from Cranberry Junction at Highway 37 north and are installing new fibre along HW 37 from the Telus fibre along HW 16 near Kitwanga to Cranberry Junction to connect to the NTL fibre. An interconnection at BC Hydro’s TCT substation will allow communications to the Treaty OPC over the fibre on the 287 kV KSM transmission line. The PFS budget also allows for the cost of new fibre plowed into the shoulder of HW 37 from Cranberry Junction to Treaty Creek, by Tahltan Communications, to provide dedicated (dark) fibre to the KSM site as may be required for mine operations in order to support remote tele-remote control and similar functions. Installation of fibre-optic cable to site will also allow for the installation of dedicated cellular service.

 

A plant wide fibre optic communication system will be installed in conjunction with the power distribution system in both the Treaty OPC and the Mine Site. A fibre-optic cable has been included in the MTT to provide communications between Treaty OPC and the Mine Site.

 

An ultra-high frequency (UHF) radio system will be used for mobile communications in both the PTMA and the Mine Site. Base stations and repeaters will be installed as necessary on ground and inside tunnels.

 

Treaty OPC wired telephone service will be provided by a Voice over Internet Protocol (VoIP) system. A local cell phone system is also planned, as is satellite television for the camps.

 

In addition, uninterruptible power supplies will be used to provide backup power to communication systems and critical control systems to facilitate orderly shutdown of process equipment and to back up computers and control systems.

 

18.9Fresh and Potable Water Supply

 

Fresh and potable water for the Treaty OPC will be supplied from nearby wells to an elevated storage tank approximately 12 m in diameter and 9 m in height. Fresh water will be used primarily as:

 

fire water for emergencies

 

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cooling water for mill motors, mill lubrication systems and reagent preparation

 

the potable water supply.

 

18.10Power Supply and Primary Distribution

 

Power generation and transmission utilities in BC are regulated by the British Columbia Utilities Commission (BCUC), acting under the Utilities Commission Act. BC Hydro generates the majority of power in BC, although there are an increasing number of private Independent Power Producers (IPPs). BC Hydro owns and operates the major transmission and distribution system in BC and is the electric utility that would serve the KSM Mine via the recently constructed NTL.

 

The utility interconnection capital cost for the KSM Mine is as set out in the facilities agreement, an arrangement approved by the BCUC in January 1991, pursuant to Order G-4-91 that sets out the rights and obligations of BC Hydro and the customer for construction, ownership, and operation of the facilities necessary for electric service. As of Q1 of 2022, a facilities agreement is in place with BC Hydro covering the required capital contributions and bonding to provide Utility construction power. A second Facilities Study and consequent Facilities Agreement for operating power will be finalized within adequate time limits for BC Hydro to carry out additional system reinforcement to carry the full 245 MW operating load. The associated costs will be covered by bonding and no further cash contributions are required.

 

18.10.1Northwest Transmission Line

 

The 344 km long, 287 kV NTL runs from the Skeena substation near Terrace, BC, to a new substation near Bob Quinn Lake (Figure 18.15). This new BC Hydro transmission line was commissioned in the summer of 2014 and currently serves the AltaGas Forrest Kerr Hydroelectric Facility and the Red Chris Mine. A tap from this transmission line will service the KSM Mine.

 

The previous BC Hydro Tariff Supplement TS37, which was enacted to recover the cost overrun of the NTL transmission line construction, has been eliminated. Thus, a payment of $254,800,000 to be repaid over 5 years after start of commercial production is no longer required. This represents a very significant saving in sustaining capital over previous studies.

 

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Figure 18.15NTL Route Map

 

Source: BC Hydro

 

18.10.2Treaty Creek Switching Station

 

BC Hydro is responsible to deliver power from the transmission system to a customer at the point of interconnection (POI). The customer is responsible to bring power from the POI to their site. For KSM, the POI will be the Treaty Creek Terminal Station (TCT). The KSM site will take electrical service via a 30 km long, 287 kV line extension from the Treaty Creek Switching Station, to be located on the NTL adjacent to Highway 37, approximately 18 km south of Bell 2 Lodge. This installation will also be in the vicinity of the Treaty Creek Access Road junction with Highway 37. Metering will also be located at this point. The TCT substation will form part of the BC Hydro system and will be constructed, owned, and operated by BC Hydro. BC Hydro has completed site selection, and is doing basic design engineering and construction has started.

 

Seabridge previously reimbursed BC Hydro for the cost of installing transmission line dead end structures when the NTL was constructed, as required to facilitate the connection of the of the proposed TCT into the grid.

 

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The power supply facility infrastructure currently under construction includes BC Hydro System Reinforcement and the Basic Transmission Extension, which is the circuit breaker and metering at the POI. It is not the transmission line to the KSM site. Construction of the TCT substation, as per the BC Hydro Facilities Agreement now in place, required a direct cash payment from KSM to cover a large part of the cost of the installation that is classified as the Basic Transmission Extension. This cost is included in the KSM Substantial Start budget. The remaining station cost is classified as System Reinforcement and is not a capital cost. However, Tariff Supplement No. 6 (TS6) Clause 5 (c) “Offset” requires that the customer provide bonding for typically up to seven years, such that BC Hydro is assured of receiving enough revenue from the KSM operation to justify the capital expenditure. Security is required in the form specified in TS 6 Clause 13 and in the amount as per Clause 5(b). The foregoing bonding and charges are set out under the tariffs and are not negotiable. BC Hydro will return part of the security each year as the offset, based on power billing, reduces the required bonding. The amount of the bonding is not included under the direct capital cost budget, but is otherwise accounted for in the PFS economics. An important point to note is that as per the tariffs, the customer must pay the actual final cost of construction, not the amount estimated by BC Hydro in a Facilities Agreement.

 

BC Hydro is responsible for obtaining all approvals and permits for the TCT substation. A formal environmental assessment of TCT under BC’s environmental assessment process for reviewing major projects was not required.

 

The Sept. 2018 System Impact Study (SIS) by BC Hydro provides for a site maximum demand of 245 MW based on 3 power supply queue positions, which were a result of two increases in the mine power supply based on two updates to the SIS, as the planned KSM size and load grew from the initial submission. As it stands, KSM have 245 MW reserved for their use. The 245 MW contract demand is not exceeded until Year 3 of production which provides adequate time for an application to increase the contract demand or install the budgeted combustion turbine.

 

18.10.3Transmission Line Extension To KSM

 

The voltage selection for the proposed 287 kV transmission line extension for KSM was based on the 287 kV transmission voltage selected for the NTL. A review of the technical requirements to serve a load in the range of 245 MW confirmed that stepping the voltage down to a lower level is not technically acceptable nor economic.

 

The KSM Mine will be responsible for the construction and operation of the transmission line extension, in accordance with the established BC Hydro tariff requirements. Line construction will utilize steel monopoles, such that the line can be generally run in the TCAR right-of-way, beside the road, thus largely eliminating the requirement for a separate access route.

 

The 287 kV transmission line from the BC Hydro TCT will cross Highway 37 and the Bell-Irving River, then closely follow TCAR along the north side of Treaty Creek for approximately 12 km to a deviation point where the line transitions from following the TCAR, to following the South Diversion Cut-off Ditch, up to main substation FLT1 at the Treaty OPC. Steel monopoles are ideal for use where a transmission line is to be constructed next to a road and in areas of high snow fall. To protect against avalanche damage, several structures will be mounted on concrete piers, to raise the pole bases above the avalanche flow (in run-out areas).

 

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The environmental assessment for the 30 km section of transmission line from TCT to Substation No. 1 was included in the KSM approved EA; therefore, approval is in place. Land tenure (a Licence of Occupation) has been obtained for the transmission line right-of-way, from the Treaty Creek Switching Station to Substation FLT1 at the Treaty OPC.

 

The 287 kV KSM transmission line includes a fibre optic cable connection to the BC Hydro NTL fibre-optic cable system, as required by the utility for system protection. This fibre connection will also carry the general communications to site for the permanent operations phase.

 

The transmission line land tenure has been obtained and clearing started. The cost of right-of-way clearing is included in the road clearing budget.

 

The 287 kV transmission line will be constructed as part of the KSM’s early works and consequently is not in the PFS budget.

 

18.10.4System Studies

 

BC Hydro performs studies to determine the cost, method, and timing of transmission system customer interconnections. Seabridge first commissioned a BC Hydro SIS for KSM in 2009 to confirm the technical viability of the interconnection. Subsequently, several updates were commissioned. The latest system impact study was completed by BC Hydro in September 2018 for a contract demand of 245 MW. A Facilities Agreement was signed in February, 2022 for the supply of 25 MW of construction power. A planned second Facilities Study will be carried out by BC Hydro to define the costs for the supply of 245MW for full operations power, as confirmed available by SIS. This study will confirm the system reinforcement bonding required for operations power. No further capital contributions are anticipated to be required.

 

System load flow studies have been performed by PFS consultants using system analysis software to confirm process plant and mine power system voltage control from no load to full load. System voltage stabilization is based on switched reactors to control light load over voltages due to 287 kV transmission line and 138 kV cable capacitance, and also assumes power transformers have automatic tap changers and that there is automatic control of the process plant synchronous ball mill drive motor excitation systems for instantaneous voltage control, as requested by BC Hydro. The mine main substation (FLT1) also includes automatic power factor correction capacitor banks to maintain unity power factor at TCT, as required by BC Hydro. BC Hydro will also include a STATCOM (solid state power electronic compensation equipment) in their TCT substation to further stabilize voltage.

 

Service from the Skeena Substation to the KSM Site will be delivered over the existing NTL single-circuit 287 kV transmission line which entered service in 2014. Occasional service interruptions and planned maintenance outages can be expected and are considered normal for mining projects.

 

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18.10.5Electric Utility Requirements, Tariffs, and Cost of Electric Power

 

The electric service to the KSM Site (including all terms and conditions such as rates and metering requirements, connection charges, and many aspects of the KSM connecting transmission line) will be in accordance with the latest edition of BC Hydro Electric Tariffs, in particular:

 

Rate Schedule 1823 – Transmission Service – Stepped Rate

 

Rate Schedule 1901 – Deferral Account Rate Rider,

 

BC Hydro Electric Tariff Supplement No. 5 (TS5) Agreement for Customers Taking Electricity under 1821 (1821 is now 1823) (TS5 is a template for the Electricity Supply Agreement with the format set as per the tariffs and is not subject to change)

 

BC Hydro Electric TS6 Agreement for Transmission Service Customers (TS6 is a fill in the blanks template for the Facilities Agreement with the format set as per the tariffs and is not subject to change)

 

BC Hydro Electric Tariff Supplement No. 74 (TS74) Customer Baseline Load Determination Guidelines.

 

BC Hydro Rate Schedule 1823 is a two-tier schedule, nominally with 90% of the Customer Baseline Load charged at economical Tier 1 energy rates, and the last 10%, plus all power above the Customer Baseline Load, charged at costly Tier 2 rates. This system is designed to encourage energy conservation, as consumption reductions due to energy conservation measures are applied against costly Tier 2 power. BC Hydro, under their Power Smart program for demand side load control, offer incentives to transmission customers to reduce energy consumption, and for new customers incentives are given for energy-efficient plant design.

 

The calculated power cost as estimated for the 2022 PFS is somewhat below regular rates due to a large reduction or elimination of costly Tier 2 energy in accordance with an efficient plant design as accepted by BC Hydro’s “Power Smart” program based on a study approved by BC Hydro. Thus, HPGR energy savings have an impact far greater than just the energy savings in the grinding area. A separate report to BC Hydro has confirmed that the use of HPGRs for KSM qualifies for these incentives.

 

BC Hydro Tariff Supplement TS6 (as approved by the BC Utilities Commission) currently requires potentially large non-refundable 500 kV transmission and generation system reinforcement charges for operations with a Contract Demand of over 150 MVA, if they require reinforcement to the 500 kV system supplying the Skeena Substation, the power source of the NTL transmission line feeding KSM. The terms covering 500 kV system upgrades could potentially apply to KSM and the required capital contribution would apply to the entire load as per the wording of the tariffs, not just the load that exceeds 245 MW The Contract Demand (peak load) for KSM is currently estimated to be above 245 MW after production Year 3, considering the large load due to trolley assist in the open pit, even considering that energy conservation and recovery measures will be implemented. Seabridge currently has an approved SIS and an agreement with BC Hydro for electric power supply with a maximum Contract Demand of 245 MW. A combustion turbine has been allotted for in the estimates after Year 3 due to trolley assist, which will cause the maximum demand to exceed the 245 kW limit.

 

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The cost of power for KSM, delivered to the 25 kV bus bars of the Treaty OPC, has been estimated as Cdn $0.0596 per kWh, based on rates effective in Q1 of 2022) including applicable taxes and energy cost savings due to BC Hydro’s Power Smart program. The KSM power cost includes the transmission line losses from the metering point at the Treaty Creek Switching Station, plus Substations No. 1 and No. 2 transformer losses and peaking power cost.

 

The KSM power cost calculation takes into account reduced rates due to BC Hydro Demand Side Management (DSM) and associated Power Smart initiatives for energy conservation measures designed into new plants (such as using HPGR grinding in lieu of SAG milling). Such measures, as may be certified by BC Hydro, serve to reduce the standard 10% of energy under the two-tier 1823 Rate Schedule that would fall under the costlier Tier 2 category. If HPGR grinding and similar energy conservation measures were not to be implemented, there would not only be greater energy consumption, but the cost of electric power per kilowatt hour for the entire KSM operation would increase as high cost Tier 2 energy would supply 10% of the load, based on the current (2022) BC Hydro rate schedule 1823.

 

Each year on April 1 (the start of their fiscal year), BC Hydro sets new rates that are applied in accordance with the tariffs, subject to BCUC approval.

 

18.10.6Treaty Plant Main Substation (FLT1)

 

The KSM 287 kV step-down Substation FLT1 will be located at the Treaty OPC and will be constructed and owned by Seabridge in accordance with BC Hydro policy, which is also the most economical solution. This substation is a critical installation for KSM. The substation equipment has been sized based on the latest PFS load list. Redundant transformer capacity was included in the design. The substation will be a GIS switchgear design, utilizing 138 kV and 287 kV gas insulated circuit breakers and bus bars, allowing a compact design contained in a building adjacent to the Treaty process plant. The circuit breakers will use point-on-wave switching as required by BC Hydro. Connections to transformers will use high-voltage solid dielectric cables.

 

The gas insulated substation will include:

 

Four transformers, each of the three winding type oil filled 75/100/125 MVA, ONAN/ONAF1/ONAF2 step down power transformers, with automatic on-line tap changers

 

nine 287 kV GIS circuit breakers

 

eleven 138 kV GIS circuit breakers

 

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one 287 kV switched reactor for compensation of the incoming 287 kV line, to limit Ferranti effect over voltages

 

four 138 kV switched reactors at the FLT1 end of the 24 km long, 138 kV cable to compensate for cable capacitance, thus controlling bus voltage

 

25 kV grounding transformers and resistors.

 

The four transformers included in Substation FLT1 will have FR3, less flammable insulating fluid and will be installed in a concrete building. They will provide redundancy, allowing one transformer to be out of service at maximum plant load. Shipping restrictions to the site were taken into account when sizing the transformers. The 138 kV tertiary windings will be connected to the 24 km long tunnel cables feeding the Mitchell (open pit mine) area Substation FLT2.

 

Substation FLT1 will also include four line-ups of 25 kV GIS switchgear, including all 25 kV circuit breakers required for power distribution to the process plant and around the Treaty OPC. The secondary distribution voltage for the Treaty OPC will be 24.9 kV, 3 phase, 3 wire, high resistance grounded.

 

Substation FLT1 does not include harmonic filters. If these are required by harmonic generating plant loads as indicated by studies during the detailed plant design, they would be located at the process plant near the harmonic sources and at the MTT train rectifier stations. As the rail system is now regenerative, local harmonic filers will most probably be required.

 

As part of Substantial Start, the gas insulated (GIS) substation FLT1 building and approximately one third of the electrical equipment is being purchased to provide construction power. The remaining equipment is included in the PFS budget for purchase, installation, and commissioning prior to production.

 

18.10.7138 kV Cable

 

Substation FLT1 will be interconnected with Substation FLT2 by two sets of three, 138 kV, single-core, 300 mm2 , stranded copper, cross-linked polyethylene (XLPE) solid dielectric insulated power cables suspended from the tunnel back in the MTT that will run between the two plant sites. The 300 mm2 (600 kcmils) conductor size quoted is the minimum physical size that vendors typically manufacture at 138 kV (due to the electric field gradient at the conductor). This is a more than adequate capacity to carry any anticipated load, including allowance for mine trolley assist, the cable charging current, etc. To limit induced sheath currents, the cable sheaths will be “cross bonded”, which is the normal design for high-voltage, high-current, single-core cable installations. Adequate (significant) space must be allowed in one of the train tunnels for these cables.

 

18.10.8Mitchell Substation FLT2

 

The 138 to 69 kV - 25 kV Substation FLT2 is also critical infrastructure for KSM. As an alternative to a standard 138 kV air-insulated outdoor substation, substation FLT2 is planned to be a GIS installation. This is a very compact design, requiring only a fraction of the space of a conventional air insulated high voltage substation and allows for the total installation to be included in a reinforced concrete building that provides a high degree of protection against geo-hazards such as avalanches and eliminates problems due to high snow fall. It also eliminates hazardous high-voltage overhead lines in the vicinity of the Mitchell OPC and requires much less plant area. The substation includes:

 

three 138 - 69 - 25 kV, 55/73/90 MVA ONAN/ONAF1/ONAF2, oil filled power transformers with automatic on-line tap changers (two, 3 phase units are provided for redundant capacity, with space provided for a third unit to cater to future load growth)

 

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eleven 138 kV GIS circuit breakers and associated bus work

 

four switched 138 kV reactors to compensate 138 kV cable capacitance and control voltage

 

ten 69 kV GIS circuit breakers connecting to site 69 kV power distribution system to the large pit trolley assist load and the facilities more remote from Substation FLT2

 

grounding transformers and resistors

 

three line-ups of 25 kV GIS switchgear for site local power distribution including 25 kV circuit breakers for all local power distribution.

 

18.10.9Site Power Distribution

 

Site power distribution in the mine area from Substation No. 2 will be by 25 kV cables and overhead pole lines locally and by 69 kV overhead pole lines to feed large trolley assist loads at more distant facilities where modular substations will step the 69 kV down to the local distribution voltage. The relatively long distances and high initial future pumping and trolley assist loads require 69 kV distribution to transmit the power and limit voltage drop.

 

18.10.10Mine Power

 

Power to the Mitchell open pit itself will be provided by 69 kV feeders and local 25 kV overhead distribution lines. The required pit 25-7.2 kV portable substations (also serving as pit switch-houses), and trailing cables for the 7.2 kV pit mobile electric shovels and drills, are included in the electrical power budget. 7.2 kV to 600 V portable substations are also included for pit dewatering. Similar installations are included in sustaining capital for the Sulphurets open pits. The open pit design also includes 69 KV transmission lines, step-down transformers, DC substations and trolley lines for open pit trolley assist.

 

18.10.11Trolley assist

 

Trolley assist for the mine haul trucks has been included in mine design. This both reduces greenhouse gas emissions and also significantly reduces fuel costs as electric power is being substituted for diesel fuel. The cost of power for the 2022 PFS is just under 6 cents per kilowatt hour and for fuel switching, BC Hydro has a new tariff (Rate Schedule 1895 – Transmission Service - Fuel Switching) that offers a reduction in electric power cost of up to 20 percent. This has recently been applied in BC for a trolley assist project at another mine site. No capital credit for Rate Schedule 1895 has been assumed in this 2022 PFS. Ninety-eight per cent of the power generated by BC Hydro is from clean, renewable resources, mostly being hydro power, thus greenhouse gas reduction is very significant when on trolley.

 

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As many large haul trucks are already diesel electric, like locomotives, rather than mechanical drive, modification by the manufacturers to adapt these trucks for trolley assist is relatively easy and several major manufacturers have proven solutions in operation, including currently in BC.

 

The mine plan has different trolley configurations over the mine life and thus savings will vary from year to year. In addition, the savings will vary to some extent based on the specific equipment purchased. However, it is fair to say on an 8% ramp there would be a fuel saving of over 95% (as the engine is on idle) and a speed gain in the range of 90%, depending on truck specifics.

 

To facilitate the high power demand of trolley assist, in the range of 35 to 40 MW, power distribution from Mitchell substation FLT2 will be via 69 kV pole lines with unit substations provided to step the voltage down to 25 kV for local distribution and utilization. The trolley system itself would include DC traction (rectifier) substations that supply power to the catenary system that consists of the catenary support aerial cables, the contact conductors and parallel heavy duty aerial feeder cables. The aerial conductors would be supported by steel cantilever cross arms and galvanized steel poles bolted to precast concrete foundations. The system is designed for relatively rapid installation and with the ability to move and re-use the lines.

 

Each truck would be fitted with a pantograph to receive external DC electric power from the catenary and the necessary power electronics to convert the DC power to variable frequency alternating current (AC) for the truck wheel motors.

 

18.10.12Construction and Standby Power

 

Modular diesel generator sets will be provided to supply construction power for tunnel driving, camps, temporary water treatment plants, plant construction sites, and other initial construction-related facilities that are not adjacent to main substation FLT1 which will supply construction power from BC Hydro for facilities in the area of the Treaty OPC. The capital and operating costs of these facilities plus local distribution including step-down transformers and overhead pole lines have been included in construction indirect costs. Fuel and operating costs for construction power are also accounted for in the construction indirect costs. The power distribution costs for supply and installation of cable and electric panel boards within the various tunnels are included in tunnelling costs.

 

Any additional costs for moving equipment and fuel to site during the early stages of the KSM construction, such as by helicopter, are included elsewhere in the capital cost estimate and are not in the construction power budget.

 

The construction generating stations are modular, complete with switchgear, and designed for PLC automatic unattended operation. Environmentally approved double-walled fuel storage tanks and associated piping are included for each power station.

 

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Several of the construction gensets will be retained after initial construction is complete and reconfigured to serve as future standby/emergency generation for the mine, process plant, and accommodation centres. The cost to refurbish construction gensets and reconnect this equipment for standby service in the permanent plant was included in the general site maintenance operating costs.

 

The estimates include the purchase rather than rental of construction gensets. The relatively very long KSM construction period will make construction genset rental uneconomic.

 

18.10.13Energy Recovery and Self Generation

 

There are several opportunities for energy recovery from process flows, as well as power generation from mini-hydro plants, taking advantage of water flows that must otherwise be diverted around the mining operations. As these energy recovery and mini hydro schemes, to a large extent, make use of facilities otherwise required for KSM, they are generally economically attractive based on the current two tier utility rate schedule and will reduce the total energy consumption of the KSM operation. The value of the generated power would be at the BC Hydro rate schedule 1823 Tier 2 energy (set at BC Hydro’s marginal cost of generation).

 

All of the listed energy recovery plants will be located within the KSM mining lease. The energy recovery plants recover energy from process plant flows.

 

All of the generating plants, similar to small IPP hydroelectric plants, will operate unattended and will be automatically controlled by PLC systems. The locally generated power will be fed into the 25 kV mine distribution power lines.

 

The generation facilities included in this PFS are summarized in Table 18.8.

 

Table 18.8Mini Hydro and Energy Recovery Power Generation

 

Name Type Installed
Capacity
(kW)
Net Annual
Generation
(kWh)
Machines
WTP* Energy Recovery 9,000 19,866,000 Pelton Turbines
Tailings Energy Recovery 1,194 10,414,000 Pumps as Turbines
McTagg Diversion* Mini Hydro 10,000 29,842,000 Francis Turbine

 Notes: * Operation continues after mine reclamation.

 

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18.11Treaty OPC and Mine Site Secondary Electrical Power Distribution and Utilization

 

18.11.1Mine and Plant Power Consumption

 

The total initial mine and process plant annual energy consumption is estimated to be 1,726 GWh based on the (HPGR) load list for the initial 130,000 t/d operation valid on average for Years 1 and 2. This equates to an average annual load of 197 MW. With a load factor (LF) in the range of 0.87 as typical for a large mine such as KSM, the annual worst month peak load (15-minute demand) is estimated as 226 MW. For Year 1 the peak loads are well below the 245 MW contract demand.

 

The corresponding figures for Year 3 with a throughput of 195,000 tpd is an annual energy consumption 2,152 GWh, an annual average load of 246 MW, and a maximum peak load of 283 MW.

 

The required utility supply will be reduced in the summer and fall by self-generation from energy recovery and mini-hydro plants. During the winter low stream flow conditions, the average self-generation will be almost zero. To prevent the KSM operational demand from exceeding the 245 MW trigger point for generation reinforcement, a proposed 22 MW peaking combustion turbine, located in at the Treaty OPC, will be operated. In addition, it would be possible to institute demand control for the large pit trolley assist electrical load. As the mine develops, there is ample opportunity to investigate the cost of increasing the BC Hydro contract demand to 300 MW. As this additional power would not be required until Year 3 of production.

 

18.11.2Power Distribution – Treaty Plant Main Substation FLT1

 

For a discussion of power distribution refer to Section 18.10.6 Treaty Plant Main Substation FLT1.

 

Ball Mills

 

The ball mills as planned were integrated into BC Hydro’s SIS and any changes to these drives will require a new BC Hydro SIS and potentially a new Facilities study.

 

The Treaty process plant ball mills are major power consumers. Each of the four ball mills (rated 14,000 kW each) in Year 1 and 2 additional mills in Year 3, will be fed via dedicated 25 kV feeders and step-down transformers to 13.8 kV. The mills will each be equipped with two, 10,000 hp, low-speed “Quadratorque” fixed speed synchronous motors, directly driving mill dual pinions via air clutches, as has been used in the industry for many years.

 

Step-down to 4.16 kV

 

The ball mills will be fed at 13.8 kV. Other large fixed speed motors (generally those rated 250 hp and greater) and large variable speed drives (generally those rated over 400 hp) will be fed at 24.9 kV or 4,160 V. As the VFDs include integral step-down transformers the larger VFDs will be fed at 25 kV, to save transformer costs. The 4,160 V supply will be derived from 25 kV to 4,160 V outdoor liquid filled step-down transformers. Redundancy will be provided by utilizing sets of two transformers, each feeding a 4,160 V metal clad switchgear line-up with the two line-ups connected by a tie breaker that may be closed if one of the transformers fails or must be taken out of service.

 

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Step-down to 600 V

 

Motors and other loads below 250 hp will be fed from one of several 600 V systems. Generally, these systems will consist of FR3 liquid insulated 25 kV to 600 V step-down transformers, feeding two line-ups of 600 V power distribution centres (with tie breaker), which in turn feed a series of 600 V motor control centres (MCCs). General power and lighting will also be fed from the 600 V system.

 

Remote Loads

 

Remote Treaty OPC loads will be served by 25 kV overhead lines. Examples of remote loads include, fresh water pumping, the TMF return water and seepage pumps and ancillary buildings.

 

18.11.3Mitchell Substation FLT2

 

For a discussion of power distribution refer to Section 18.10.8, Mitchell Substation FLT2.

 

Power Feed to Pits, Trolley Assist and Primary Crusher

 

The Mitchell primary crusher will be fed from substation FLT2 by a 25 kV cable. The Mitchell pit 25 kV overhead power line will be fed from a section cable leading into the substation.

 

The mining electric shovels and drills will be served at 7.2 kV via portable 25 to 7.2 kV step-down substations fed from the perimeter pit pole line. The estimates include appropriate lengths of trailing cable and couplers. 7.2 kV to 600 V portable step-down substations and trailing cables are also included for pit dewatering.

 

A 69 kV GIS circuit breaker and cable will feed an overhead pole line supplying remote loads including the truck shop, WTP, Selenium WTP, WSD pumping, explosives facility, permanent camp, and also connecting to the mini hydro and energy recovery power plants. There will be local unit substations stepping down from 69 kV to the local distribution voltage.

 

Because of the large power demand associated with pit trolley assist, 69kV distribution lines and 69kV unit substations are included in the trolley assist power supply to step the voltage down to 25kV to supply trolley assist DC substations.

 

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18.12Permanent and Construction Access Roads

 

The Mine Site and PTMA construction will require a combination of permanent and temporary access roads. Preliminary design work was performed pre-2009 utilizing remote sensing data (LiDAR) and was ground-truthed between 2009 and 2012. Primary access to the mine during construction will be via the CCAR and TCAR, with a winter access road being utilized during the initial construction phase. In total, seven roads are proposed with varying design criteria that match that anticipated traffic volumes generated at the Mine Site and PTMA during construction and operation. A map outlining the proposed roads is shown in Figure 18.16.

 

Current proposed permanent access roads include the existing 59 km long resource access route from Highway 37 to the former Eskay Creek Mine and camp facilities. The proposed 33 km long CCAR will commence near the southern limit of this existing road, and extend south then west to the proposed Mine Site.

 

The TCAR network provides access to the Treaty OPC, the TMF, and the MTT Saddle Area. It will include a 30 km two-lane access route from Highway 37 to the Treaty OPC, TMF, and Treaty MTT portal, and include portions of the TCAR and NTAR.

 

The current proposed access roads include the:

 

Eskay Creek Mine Route

 

CCAR

 

TCAR

 

Lower NTAR (early mine life)

 

Upper NTAR (early and mid-mine life)

 

Cut-off Ditch Access Road

 

TMF Service Roads

 

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Figure 18.16Proposed Access Roads Network

 

 Source: Tetra Tech

 

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18.12.1Route Descriptions

 

The current updated route descriptions, including relocations of the road alignments, are provided in the following sections.

 

Eskay Creek Mine Access Route

 

Seabridge plans to use portions of the existing Eskay Creek Mine Access Road, linking Highway 37 to the proposed CCAR.

 

This road was constructed in 1993 to provide access to Homestake Canada Inc. Eskay Creek Mine. In 2002 the mine was assumed by Barrick Gold Corp. (Barrick) until its closure in 2008. The property is now under option with Skeena Resources Ltd. The road commences at Highway 37, south of the Bob Quinn Forest Service Road, and follows the Iskut River Valley west for approximately 38 km to the crossing of Volcano Creek. The road was originally designed as a single lane, 5 m wide gravel road, with a nominal design speed of 60 km/h. Substantial portions are built to a nominal 8 m wide (double-lane standard), providing ample passing opportunities.

 

An overview evaluation of the road condition and its suitability for Seabridge’s requirements was conducted; findings are summarized in McElhanney (2013).

 

Coulter Creek Access Road

 

The CCAR will be constructed as a single lane (6 m surface), radio-assisted road with pullouts, and will connect the existing Eskay Mine road to the KSM mine site.

 

Heading southwest from near the end of the existing Eskay Creek Mine Access road (approximately 59 km off of Highway 37), this road will follow an existing mine access road for approximately three or more kilometres towards Tom MacKay Lake. It will then descend out of the alpine meadows, along the height of land between Coulter Creek and the Unuk River. Construction of the first 8 km of road began in 2021 and is ongoing in 2022. Additional road building to the Unuk river is under consideration by Seabridge for the 2023-2024 timeframe.

 

The proposed three-span bridge crossing of the Unuk River will be 88 m in length. The Unuk River is a major crossing and will need to meet the requirements of the Navigable Waters Protection Act. Beyond the Unuk River, the route traverses a short section of low-lying wet and swampy areas and then starts to climb steeply through a series of switchbacks into the Sulphurets Valley and canyon. The alignment extends to the HDS WTP where the CCAR road design (and Special Use Permit [SUP] boundary) ends. Beyond this point, the road design is determined by the mine development, and is the responsibility of the open pit mine designer MMTS.

 

Eskay Mining Corp. owns a large claim block immediately southwest, west, and northwest of KSM. Parts of the proposed CCAR is situated on mineral tenures held by Eskay Mining. On July 5th, 2021, Seabridge and Eskay Mining Corp entered into an agreement where both companies will share the costs equally on construction and maintenance of the first 8km portion of this road.

 

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Additionally, Skeena Resources owns mineral claims and leases covering other parts of the CCAR immediately north of Eskay Mining’s exploration project. Skeena owns the previously operating Eskay Creek gold and silver mine, last operated by Barrick, and is developing new resources elsewhere on its claims. Seabridge and Skeena are actively discussing a cooperative agreement towards mutually beneficial development of the access road.

 

Treaty Creek Access Road (HWY 37 to KM 17.9)

 

The TCAR will leave Highway 37 approximately 19 km south of Bell 2, and head west. It will be constructed as a two lane (8 m surface) all-season road to the junction of the Treaty Creek and North Treaty Upper road at km 17.9.

 

Meetings were held between McElhanney, Seabridge, and the provincial Ministry of Transportation and Infrastructure (MOTI) to discuss and establish a set of design criteria for the proposed intersection at the Highway 37/TCAR location. Construction design work has been completed and the intersections’ construction is fully permitted by MOTI.

 

Initially, the TCAR will follow a former forestry access road. A three-span 119 m long bridge is proposed for the crossing of the Bell-Irving River. This is a major river and will need to meet the requirements of the Navigable Waters Protection Act. Construction of this bridge began in Q1 2022 and will be commissioned in Q3 2022. The proposed road follows the north side of the Treaty Creek Valley and construction began in 2022. It will generally be located between the flatter riparian zone below and the steeper avalanche-prone terrain on the north slope. The TCAR will continue as a double-lane road further west to a future intersection with NTAR. Heading west from there, the TCAR will transition into a single-lane road leading to the MTT saddle area.

 

Treaty Creek Access Road (KM 17.9 To Tunnel Saddle Access Portal)

 

Beyond km 17.9 road intersection and heading west up the Treaty Creek Valley, the TCAR will provide construction period access.

 

North Treaty Creek Access Roads

 

There are currently three permanent access road alignments proposed within the North Treaty/Teigen Creek valley. They are referred to as the lower NTAR, the upper NTAR, and the Cut-off Ditch Access Road.

 

Lower NTAR

 

Earlier access can be obtained by constructing the lower NTAR. This will leave the TCAR at approximately km 16.9 and follow the lower valley. The lower NTAR will be quicker to build. It will result in a slightly shorter travel distance between Highway 37 and the Treaty OPC and TMF, and with generally flatter grades. This road would be used for approximately the first half of the mine life, until such time as it is necessary to construct the southeast tailings dam. The primary purpose of the NTAR is to provide early access to the lower valley for the construction of the tailings dam(s). Eventually, the north section of this road would be buried by the southeast tailings dam.

 

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Upper NTAR

 

The upper NTAR will leave the TCAR at approximately km 17.9. It will traverse approximately 12 km north from the TCAR to the Treaty OPC and TMF. The road would then parallel the proposed drainage cut-off ditch, which will divert drainage off the west slope of the valley, north to the Teigen Creek Valley.

 

Cut-off Ditch Access Road

 

The proposed cut-off ditch access road is to provide construction and maintenance access only. The power transmission line to the process plant will follow this route.

 

North Treaty/Teigen TMF Service Roads

 

Access to the Tailings Management Facility will be provided by approximately 28.4 km of 6 m wide service roads (with pullouts). These roads will provide access to the east side of the North Treaty and Teigen Creek valleys, including a water well. The TMF service roads include the following segments:

 

South Teigen Road 11 (5.5 km)

 

South Teigen Road 11a (3.15 km)

 

South Teigen Road 12 (9.1 km)

 

South Teigen Road 12a (0.3 km)

 

Upper Dam Road 12b (1.2 km)

 

South Teigen Road 15 (3.7 km)

 

South Teigen Road 15a (2.7 km)

 

Water Well Access Road (2.7 km).

 

18.12.2Road Design Requirements

 

The KSM Site access roads are classified as resource development roads.

 

The Eskay Creek Mine Road and CCAR will be maintained for the life of the mine to support the mine development, transport of oversize loads, and to provide alternate emergency access. However, these roads will only be used seasonally, and not used during winter months.

 

The CCAR will be a single-lane (6 m surface) radio-assisted road with turnouts and widenings to allow the largest vehicles and loads access to the Mine Site. The CCAR would have some sections with sustained maximum grades of 12%. Design speeds vary greatly, in large part controlled by the terrain.

 

The proposed TCAR to km 17.9, and the connecting upper and lower NTARs, will be required for permanent access to the Treaty OPC and TMF, and to the Mine Site via the MTT tunnels. These will be two-lane roads (8 m finished surface), capable of carrying highway legal axle loading year-round. The roads will provide access for supplies, equipment, and crew transport, and be used for hauling concentrate to Highway 37.

 

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Alignment controls such as maximum 10% sustained grades (11% short pitch), and minimum 100+ m radius horizontal curves are utilized for the higher-traffic volumes anticipated on this route. Appropriate vertical profile crest and sag curve “K” values are applied. Except for a few control sections, the nominal minimum design speeds for these sections of road is 50 km/h.

 

All bridges will be designed to BC Forest Service minimum L100 loading (90,680 kg gross vehicle weight [GVW]) and minimum 1.5 m clearance above the estimated 100-year flood level (Q100). Select structures must meet additional requirements, as prescribed by the Navigable Waters Act. All bridges, including those on the TCAR, will be single lane. The exception is the Bell Irving River Bridge that was designed to a L150 loading.

 

Major culverts have been designed to pass the estimated 100-year flood level (Q100) with no headwater.

 

18.12.3Design Progress

 

Early Design Work (Pre-2009)

 

Utilizing LiDAR survey data acquired in summer 2008 and fall 2011, and the resulting digital elevation models developed, the preferred preliminary access routes identified by Seabridge and McElhanney were prepared. The preliminary road alignments were subsequently located in the field using GPS, and marked with flagging. The objective was to locate and map the most appropriate road alignment for each route based on design standards established by the team. Road alignments and cross sections took into consideration the requirements for both construction and operation phases.

 

The routes were assessed in the field and adjusted as deemed appropriate. Often several preliminary lines were investigated in order to achieve the preferred road location. Selecting the ultimate road locations was an iterative process involving both field and office design. Based on the preliminary layout, terrain information was gathered, along with bridge and major culvert crossing information. The originally flagged centerline provided a base for follow-up environmental and geotechnical assessments.

 

Based on the field reconnaissance, design standards, and associated surveys and preliminary assessments/input by other sub-consultants; preliminary road design plans and profiles, conceptual bridge and stream crossing structure designs, and construction cost estimates were prepared. Engineering assessments were conducted in conjunction with available geotechnical and environmental studies of all proposed routes. The field reconnaissance and bridge site surveys confirmed the accuracy of the LiDAR data.

 

Design Work (2009-2012)

 

From 2009 through 2012, consultants BGC and Rescan (now ERM) conducted further geotechnical and environmental assessments, respectively, on the proposed, and altered routes. Where appropriate, McElhanney’s QP at that time accompanied these consultants in the field to make joint determinations with respect to the most appropriate locations for specific sections of road. McElhanney worked with these consultants to optimize the road locations and designs.

 

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All proposed final road locations were marked with survey flagging. Flagging was marked with survey crew and date information (black felt marker), and locations identified by real-time kinematic (RTK)-GPS survey methods. Select field station references are now indicated on the road plan/profile design drawings for cross reference.

 

Work included gathering of detailed information to be utilized in refining the design(s), including soils, vegetation, potential borrow/waste areas, drainage culvert requirements, and other relevant information. Stream crossing surveys were completed on “smaller” tributaries. Generally, this includes any stream with an estimated 100-year peak flow of 6.0 m3/s or greater. Details for all such structures were completed to satisfy the requirements of the BC Ministry of Forests, Lands and Natural Resource Operations (MFLNRO) for the SUP applications. Preliminary stream crossing structure designs have been completed for all sites requiring bridges or major culverts.

 

Design Updates (post-2012)

 

Additional field work was conducted in 2012 along the proposed access routes by McElhanney, BGC and ERM. New information was incorporated to optimize the road and structure designs. Horizontal and vertical alignments were modified to best meet the environmental, geotechnical and archaeological concerns and requirements. During 2012, work was completed to locate potential borrow and waste sites, at appropriate locations to accommodate road grade construction requirements along the access corridors. Provision was also made to identify areas of potential gravel sources for road construction and surfacing materials.

 

Log landing locations were identified for decking of timber felled during right-of-way clearing operations. Log landings were located, and spaced, as appropriate for logging/skidding operations. Timber maturities/volumes, etc. were considered in establishing proposed landing locations.

 

The proposed right-of-way (clearing) boundaries now defined on the design drawings are minimum 30 m wide, expanded to include proposed borrow and waste areas, and log landings as described above. The approved SUP boundary limits extend a minimum of 37.5 m either side of proposed road design centerline (total 75 m width), widened as required to incorporate additional areas as otherwise defined. The intention is that this will provide some flexibility in adjusting the design or construction methodology as may be required due to actual field conditions encountered, without requiring multiple amendment to the SUP during the construction period.

 

The potential for ML/ acid rock drainage (ARD) has been assessed for all access road rights-of-way. Additional assessments were conducted in late 2014 for km 0 to 6 of the CCAR which had been flagged by government agencies as an area of special concern.

 

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The SUPs for road construction associated with the site access construction were granted by the Provincial MFLNRO Road Division office on September 27, 2014. There is a requirement to provide a security, payable to MFLNRO, prior to commencing construction. Payment for SUP S25750 in the amount of $520,000 for CCAR was made in July 2021 and for SUP S25751 in the amount of $4,300,000 for TCAR in January 2022.

 

Access road surveys, designs and drawings were prepared in conformance with standards provided in the then most current version of the BC government Forest Service Engineering Manual (November 29, 2012). Detailed engineering of specific slope stability measures will be subject to review by the geo-technical engineer(s), immediately in advance of, and during construction activities.

 

Bridge and major culvert structure site plan surveys, designs and general arrangement drawings have been prepared in accordance with MFLNRO Road Division requirements and current industry standards. General arrangement design drawings have been signed and sealed independently by a professional engineer registered to practice in BC. Detailed structure design details will need to be completed in advance of construction.

 

During 2021, field survey data was collected for all major crossings on TCAR, and final design drawings were prepared for TCAR, NTAR and CCAR. Road construction drawings and accompanying specifications for CCAR were prepared by McElhanney and used in tender and initial construction on the first 8 km portion of road. Additionally, CCAR construction design was revised from km 2.2 to 5.1 to avoid an overlap with Lease No. 740715 (DL 7325) and approved by the Ministry of Forest, Lands and Natural Resource Operations and Rural Development in August 2021. Road and bridge construction drawings and accompanying specifications for TCAR and NTAR were prepared by Allnorth and used in competitive tender for major bridge and road building projects initiated in 2022.

 

18.13Logistics

 

The KSM Site is currently accessible by helicopter only. Helicopter support will be used initially to transport equipment, supplies, and personnel prior to completion of the access pioneering roads to the Mitchell valley and the Treaty plant site.

 

Once surface routes of TCAR, NTAR and CCAR are established, construction equipment, materials, personnel and consumables will be imported to various work fronts located across the KSM site through these routes and dependence on helicopter support will reduce significantly.

 

Copper concentrate will be transported from the KSM site by trucks to a deep-water port facility in Stewart, BC, and then loaded onto oceangoing vessels. Two full service ports exist at Stewart, each with roll-on/roll-off freight handling capacity and either are presently, or would by the time operations begin, be capable of concentrate storage and handling to ship loading.

 

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The port is at the head of the Portland Canal, which is a 150 km fjord that is the northernmost ice-free port in North America. The port is accessible via truck on Highway 37A; however, there is no direct rail service. Concentrates from other northern BC mines are currently shipped from this port. In addition, there is interest from other operations in the region for concentrate handling services at the port.

 

For the purposes of this PFS, Tetra Tech calculated that the copper concentrates will be shipped in bulk, and that the average annual output of copper concentrate for the initial two years, including ramp-up period in Year 1 and a lower nominal process rate of 130,000 t/d in Year 2, will be 345,000 dmt/a. After completion of process plant expansion to 195,000 t/d in Year 3, the annual copper concentrate production is expected to increase to approximately 505,000 dmt/a (Year 3 to Year 10 average).

 

Molybdenum concentrate will be transported in bags from the KSM site via trucks to the port of Prince Rupert. The bags will be transferred from the trucks to containers and then delivered to Fairview Terminal for ultimate loading onto an oceangoing vessel.

 

It was assumed that the processed molybdenum will be loaded in 1 t bags for transport purposes, and that the average annual output will be approximately 1,077 dmt/a molybdenum in the first two years of operation prior to process plant expansion to 195,000 t/d in Year 3. The estimated average annual molybdenum production during LOM is about 3,854 dmt/a.

 

18.14Preliminary Construction Execution Plan

 

The Construction Execution Plan describes how the KSM Mine could be constructed. It is a plan in the preliminary planning stage and briefly defines the construction elements required to successfully execute construction management for KSM. It is notable that the mine development approach will be dictated by the majority owner operatory of KSM when it goes into construction and may deviate from that described herein.

 

18.14.1Introduction

 

The KSM Mine will require up to six years to construct, depending on the scope of Early Works completed prior to full development approval based on a Bankable Feasibility Study. The construction scope is intended to meet the following key objectives:

 

deliver an optimized, safe, and environmentally compliant mine development in accordance with the systems and procedures in place

 

perform construction activities safely, striving for zero recordable accidents

 

construction conducted in accordance with the IBA that are in place with the First and Treaty Nations

 

ensure that regulations, license agreements, applicable specifications, and standards are met

 

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complete construction within the agreed schedule, not exceeding the budget, and delivering the full scope as described in the construction authorization.

 

18.14.2Early Works Plan

 

An Early Works Plan will be developed to ensure certain key infrastructure (e.g., construction camps) and support services (e.g., catering) are in place early during construction and functioning efficiently for a successful construction program. Portions of this early work scope are being executed by KSM starting in 2021.

 

The following planning and field construction focus areas must be addressed in the Early Works Plan:

 

Planning

 

permit review and renewal plans

 

construction procedures

 

staffing, recruiting and labour relations plan, including commitments in accordance with the IBAs that are in place with the First and Treaty Nations

 

contracting strategy and plan, including commitments in accordance with the IBAs that are in place with the First and Treaty Nations – vetted and approved by the Owner

 

site access plan – pioneer roads, bridges, followed by completed permanent roads

 

health, safety and security (HS&S) management plan and manual

 

site and camp rules and regulations plan

 

environmental and cultural sensitivity awareness training plan

 

health and hygiene program

 

site safety and security orientation program

 

geohazards and avalanche management plan

 

logistics supply and materials management plan for early material requirements, including helicopter support

 

employee transportation plan for early construction program; air and ground planning required

 

environmental management plan to manage sediment control, waste, spills, fueling, etc. and wildlife management plan for early construction activities

 

community relations plan

 

quality management system

 

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safety and emergency response plans including content related to medical facilities and medical attention, emergency medevac, etc. Final Level IV resource loaded construction schedule

 

Final development execution plan

 

Field Construction

 

establish explosive supply storage and controls

 

identifying and proving borrow pits

 

sourcing road materials and aggregates; setting up crushing and screening facilities

 

build pioneering roads along CCAR and TCAR alignments to establish road access to the Mine Site and Treaty OPC

 

establish aggregate plant, aggregate wash plant, batch plant installation and supply of cement and aggregates

 

install asphalt plant and supply of asphalt

 

develop fuel supply and storage locations on site immediately upon achievement of road access build construction camps

 

establish temporary construction power – standalone power supply systems (gensets) in containers with fuel systems; build TWTP ponds and muck pads, and install TWTP’s where tunnelling is on the critical path (e.g., MTT)

 

18.14.3Construction Scope

 

The construction scope outlined in this section, summarizes the main infrastructure items constructed as permanent facilities or activities required to support permanent constructions within and surrounding the Mine Site, Mitchell OPC, MTT and PTMA:

 

Mine Site:

 

qupgrades to the existing Eskay Creek Mine Access Road

 

qCCAR (33 km, permanent access road; begun in 2021)

 

qMitchell OPC (primary crushing at a peak of 10,000 t/h)

 

qWSF (WSD crest built to elevation 716 masl) and ancillary facilities

 

qHDS WTP with three large clarifiers and sludge storage to initially process up to 3.0 m3/s

 

qsix TWTPs and associated muck piles and treatment ponds

 

qsurface water diversion tunnels (MDT and MTDT)

 

qpower distribution comprising construction gensets, overhead transmission lines, power distribution, and substation

 

qlogging, site clearing and grubbing (overburden, soils stockpiles and large woody debris for future reclamation purposes); rough/finish grading; structural excavations and fills; foundations; steel erection; architectural, mechanical, electrical and instrumentation works

 

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qMine Site ancillary buildings and infrastructure such as camps (permanent and temporary), fuel storage yard, sludge storage building, diversion ditches and bypass pipelines, material handling, temporary truck assembly yard, energy recovery plants

 

MTT:

 

qtwo 21.9 km long tunnels plus ancillary excavations train system for ore transport that is capable of delivering 130kt/d of coarse ore to Treaty OPC

 

qtwo 15,000 t ore bins and associated transfer conveyor from the Mitchell OPC

 

qpower distribution infrastructure

 

qtrain maintenance building

 

qtrain loading/unloading facilities

 

PTMA and Saddle Area:

 

qTCAR (30 km, permanent access road including major bridges)

 

qTMF (North Dam built to 930 masl, and Splitter and Saddle dams built to 935 masl, with a fully lined and drained basin for placement of CIL tailing)

 

qCOS and transfer conveyors

 

qprocess plant built for 130,000 t/d average throughput, including concentrate storage and load out, select portions of process plant built out to accommodate Year 3 expansion to 195,000 t/d

 

qtwo TWTPs and associated muck piles and treatment ponds

 

qpower distribution comprising construction gensets, overhead transmission lines, power distribution, and substations

 

qlogging; site clearing and grubbing (overburden, soils stockpiles and large woody debris for future reclamation purposes); rough/finish grading; structural excavations and fills; foundations; steel erection; architectural, mechanical, electrical, and instrumentation works

 

qPTMA ancillary buildings and infrastructure such as camps (permanent and temporary), cold storage, fuel storage yard, diversion ditches and pipelines, and material handling

 

Infrastructure (major, both on- and off-site):

 

qconcentrate storage and loading at the Port of Stewart

 

qswitching station at TCAR turn off from Highway 37(by BC Hydro, begun in 2022)

 

q287 kV overhead power line from switching station to PTMA

 

qfibre optics along power line right-of-way and tie-ins

 

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qHighway 37 marshalling yard and Highway 37 turnoff, including site security infrastructure.

 

18.14.4Construction Schedule

 

The 2022 PFS construction schedule was compiled in accordance with the AACE® International (AACE®) recommended scheduling guidelines (level of detail is at Level 2), with a Class 4 definition. The construction schedule is estimated to be six years and has been designed to accommodate major seasonal and environmental constraints.

 

Critical path consists of pioneering roads along the alignments of the principal access arteries to the construction areas (CCAR and Upper TCAR), MTT, and the train system that will connect the Mine Site with the PTMA. Prior to completion of the remaining KSM site access pioneering roads on CCAR, helicopter support will be utilized to support early construction activities at the Mine Site. The strategy is to establish site access pioneer roads as early as possible to reduce heli-support costs. Upon completion of the access roads to full width, major equipment and materials will be transported to site via ground freight.

 

Major site infrastructure such as the WSF and the TMF could potentially be on the critical path should the MTT tunneling duration be shortened. A preliminary construction schedule has been developed with a start date for the construction program assumed for mid-Year -6. Contractors to begin construction on the CCAR and Upper TCAR construction would assume to mobilize for mid-Year -6.

 

Mine Site pioneering begins with the development of the site access roads to the major infrastructure pads such as HDS WTP area, WSF, Mitchell OPC, MTT portals, water diversion tunnels, batch plant, TWTPs, accommodation complexes and powder storage, initially from when CCAR extends into Mitchell Valley. Early works material and equipment will mobilize to the Mine Site via heavy helicopter lift then general construction materials and heavy earth moving equipment will mobilize via the CCAR. PTMA construction will utilize the TCAR/NTAR route established in Early works to transport all material and equipment.

 

The construction duration is estimated at 68 months. The summary schedule is shown in Figure 18.17.

 

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Figure 18.17Construction Schedule Summary

 

 Source: Tetra Tech, 2022

 

18.14.5Engineering and Procurement

 

Engineering and procurement activity will be managed by teams of professionals who will report up through the EPCM contractor’s directorate. The Engineering Team will provide the required drawings, specifications, and documents to the Procurement Team in order to purchase all equipment and materials for the construction, and to allow field construction of the scope to the design intent. The EPCM contractor’s scope will include process facility and infrastructure engineering, including managing specialty contractors for major dam and tunnel designs. Mine designs will be developed and delivered by the Owner’s Team.

 

The Procurement Team will receive the engineering documentation and obtain multiple quotations that meet engineering specifications and provide a purchase recommendation to EPCM director. After EPCM director approval, the Procurement Team will purchase equipment and materials and arrange all logistics to deliver the items to the construction site ready for installation. The Procurement Team will also be responsible for establishing service contracts for engineering and field construction services.

 

Both the Engineering and Procurement teams will be including commitments in accordance with the IBAs that are in place with the First and Treaty Nations.

 

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18.14.6Construction Management

 

The Construction Management Team will be responsible for the management of all activities related to the construction management scope that includes all construction activity in the mine, process, and infrastructure areas (on site and off site). Mining activity, environmental monitoring and reporting and community affairs, which will be the accountability of the Owner’s Team.

 

The Construction Management Team will oversee the installation of all materials and equipment according to engineering and manufacturers specifications and build the facilities to satisfy the design intent and be fully operable. The Construction Management Team is also accountable for construction activity and the construction site until hand over to the Owner following dry commissioning.

 

18.14.7Construction Supervision and Contractor Management

 

The objective of all site construction activities is the timely and cost-effective completion of the construction facilities in a safe manner to the design intent and required standards in accordance with schedule. Construction supervision staff, while ensuring that standards are maintained, will provide all oversight management to contractors in achieving this objective.

 

The Contracts Management Group, which falls under the responsibilities of the site procurement manager, will use an integrated data management system to track contractor invoicing, changes, and requests for information (RFIs). The EPCM contractor will develop a comprehensive set of procedures, in conjunction with and approved by the Owner. These procedures will outline the requirements for the execution of the administrative activities.

 

18.14.8Contracting Packaging and Strategy Overview

 

The preliminary construction strategy includes dividing the construction into contract packages including commitments in accordance with the IBAs that are in place with the First and Treaty Nations. During the contractor expression of interest and pre-qualifications phase and during the advancement of detailed engineering, the contract packages will be combined to reduce the total number of contracts and form a final contracting strategy for the construction.

 

18.14.9Site Organization Structure

 

A proposed EPCM site organization structure has been developed to provide a balanced combination of senior managers, area managers, engineers, superintendents, and discipline specialists, to provide the Owner and contractors continuous support during the installation period. A high-level organizational chart is provided in Figure 18.18.

 

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Figure 18.18EPCM Organizational Chart

 

Source: Tetra Tech

 

The site organization and staffing plan has been designed by work type (e.g., engineering vs. cost controls) with the geographical constraints of a large construction site incorporated.

 

Each of the two major construction sites will have a dedicated health and safety manager and multiple health and safety representatives to assist contractors with the daily issues and training requirements.

 

18.14.10Environmental and Community Affairs

 

Environmental and community affairs during construction will be managed exclusively by the Owner’s Team to maintain independency from the EPCM Team. Environmental knowledge and community relationships have been developed through historical activity at the construction site and these relationships must continue to be managed appropriately in the context of regulatory permits granted and the societal expectations that have been expressed to the Owner’s Team. These activities will continue to be of paramount importance to the Owner well beyond the construction period, thus are best addressed by the mine owning entity that will have presence throughout the mine life.

 

A cultural awareness training program will identify and provide an overview of the various Indigenous groups who have an interest in the development, focusing on their rights as it pertains to their traditional use of the natural resources of the area. Contractual obligations negotiated between the Owner and the various groups as components of Impact Benefit Agreements will also be reviewed at a very high level.

 

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18.14.11Pre-commissioning/Commissioning

 

The commissioning period starts in any specific work area after all materials and equipment have been installed to design specification and the EPCM contractor certifies installation complete and hands the area over to the commissioning team. For the purposes of this PFS update, commissioning starts after equipment or material installation for a system or work area is complete and ends when ore starts to be processed to yield a revenue stream (i.e. the battery limit between Year -1 and Year 1). In this PFS the EPCM contractor’s scope includes commissioning through dry commissioning when work areas are handed over to the Owner’s commissioning team. The Owner’s commissioning team executes wet and process commissioning with select operating staff and professionals that are separate in reporting line and accountability from the operations staff. When these phases of commissioning are complete, they will be handed over to Owner’s operations staff who will operate the facilities through ramp up in Year 1 and on to normal operation.

 

18.15Owner’s Implementation Plan

 

The KSM Mine will be constructed as outlined in Section 18.14 and in the time frames indicated in the construction schedule in Section 18.14.4. It is the Owner’s responsibility to attain, and renew when necessary, all environmental and operating permits allowing site access road development, mine construction, and all mine operations for KSM.

 

This Owner’s implementation plan described herein attempts to provide a preliminary outline to the key responsibilities and actions the Owner’s Team will take, including interaction with EPCM contractors during the construction stage and commitments in accordance with the IBAs that are in place with the First and Treaty Nations. It is assumed KSM will be developed as a JV or consortium of two or more companies that will form a partnership to build and operate the KSM Mine. A JV organization would allow KSM’s partners to reduce risk and spread capital expense. The “Owner” referenced in this section is synonymous with this JV organization. It is further assumed that the structure developed will assign decision making authority to the majority stakeholder to eliminate bureaucracy and streamline development and mine production decisions. The KSM Mine will therefore have its own operating structure and reporting line through the JV partnership, maintaining its own profit and loss accountability to the JV partners. The Owner’s organizational structure will have a KSM president with multiple reporting lines through a six-layer organization. Site based reporting lines to the president comprise construction, mine, and process with on-site administrative functional support as necessary to enable the Owner’s Team’s success. Additionally, off-site business and external relations functions would also report to the KSM president.

 

A central office in Vancouver is not anticipated. Instead, satellite offices will be located in Terrace, Smithers and Stewart, BC to facilitate support functions sufficiently close to the construction site to provide effective support. The implementation plan described in this section highlights some key tasks required for execution by the Owner’s Team over the course of construction. There are two initial critical tasks for the Owner’s Team, starting with the identification and hiring of the KSM president, who will initially select a team, who in turn will do the same for their respective teams. This process is expected to be repeated throughout the course of construction until the entire organization has been built, while directing and supporting construction in various roles.

 

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The second initial critical Owner’s Team task is the engagement of an EPCM contractor early in the development schedule to drive the majority of the scope that resides outside of the Owner’s direct responsibility. The type of contractual arrangement between the Owner and EPCM contractor has not been established, as it relies on the strategy that will be developed after forming the JV and may be influenced heavily by the operating style of the JV partners.

 

The Owner will manage any early engineering work required to prepare design documents that support permit applications or renewals and compliance reports for permits issued by the Province of British Columbia and the Government of Canada. Site road access permits, construction camps approvals, and limited site development permits have been obtained. Additional permits will be required by the KSM Mine to ensure the completion of construction and the initiation of long term operations. During construction, the Owner will be responsible for:

 

mine development/construction including pre-stripping

 

supervision of mine fleet assembly

 

all environmental baseline monitoring, permitting and compliance

 

operation of temporary water treatment and sewage treatment plants

 

community and governmental relations

 

Treaty and First Nation relations

 

competitively bidding, adjudication and award of EPCM

 

Owner’s Team recruitment

 

training of operating personnel for the Mitchell pre-mining phases

 

medical support

 

verification surveying for measurement and payment

 

on boarding all G&A staff for both on-site and off-site positions in advance of commissioning to assist as part of the Commissioning Team and to develop process and procedures for each department/function to efficiently support operations

 

all personnel required for ultimate mine and plant start-up and operations.

 

The Owner will recruit and train technical operations and administrative staff to work in the following locations:

 

Treaty OPC Operations and Water Management – process plant and train operations, maintenance, TMF operations, security, and administrative personnel

 

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Mine Site and Water Management – operations, maintenance, security and warehousing personnel

 

Smithers Office – proposed to be developed to service all of KSM’s needs for external relations comprising governmental affairs, environmental management, permitting and compliance, public and community relations and communications, First Nations and Treaty Nation relations

 

Terrace Office – proposed as a business centre where home office support will be based for these administrative functions: supply chain and logistics, human resources, IT, accounting functions, tax, business analysis, legal and audit. Health, safety and loss prevention may have occasional presence in this office, but will be primarily based on site to support ongoing operations as they develop

 

Stewart Port Site – management of deliveries and security for incoming construction equipment/materials and outgoing concentrate shipments.

 

A conceptual onboarding plan for specific G&A functions will be developed prior to turnover of constructed areas of the site from EPCM to the Owner’s Team and is programmed to be well in advance of the turnover. This will allow sufficient time for the development of internal KSM Mine processes and procedures as a means to facilitate a smooth mine start up. The early onboarding plan is intended to cover gaps in service areas that may not have been detected in this early stage of design.

 

The time sequences for key Owner activities by year are outlined in Table 18.9. Note that this task list assumes that a FS has been completed or is running concurrently with the construction start and that either full or conditional/partial construction funding will be granted by the Owner ahead of the construction start.

 

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Table 18.9Owner’s Key Activities by Year

 

Year Activities
(Year -6)

 complete FS by Q2 Year -6

 renew early works permits, as necessary, to support initial construction

 recruit KSM president and project directors

 recruit and onboard key department leads for process, human resources, business manager, and environment and community relations

 initiate training programs required to meet First and Treaty Nation employment objectives 

 establish a business office in Terrace, BC 

 implement site environmental monitoring program at the start of construction 

 recruit and install at site environmental monitors, TWTP and sewage plant operators, field coordinators to support early works construction mainly for road building and camp construction 

 competitively bid, adjudicate, and award EPCM services 

 in conjunction with the EPCM contractor, develop a detailed execution plan for the construction leveraging all previous engineering, construction planning, and environmental permitting work. 

 establish project governance between Owner and EPCM teams 

 finalize detailed engineering work for early phase construction activities 

 release early works contracts for remaining road and bridge construction (CCAR & Upper TCAR; critical path) 

 initiate Treaty Creek fish habitat compensation construction 

(Year -5)

 manage EPCM contractor focusing on detail engineering and long-lead procurement activities off site, and early works construction on site 

 augment site Owner’s Team adding environmental services, survey (measurement and payment), medical services 

 expand Smithers, BC office to accommodate larger office headcount 

 initiate Mine Site development 

 establish road to Mitchell P8 quarry to support WSD construction 

(Year -4)

 augment site Owner’s Team in the following areas: process operations, site administration, security, site project management, human resources and environmental services; process operations staff are expected initially to reside in the EPCM contractor’s office to guide detailed design and process equipment selection 

 continue to manage EPCM contractor focusing on detail engineering and procurement activities off site and early works construction on site; procurement focusing on large equipment purchases required at site in Year -3 and beyond 

 continue Mine Site development 

 start delivery P8 Quarry rock to WSD for construction 

(Year -3)

 complete final detailed design for all remaining project scope and finalize all equipment purchases. 

 initiate enterprise computer systems set up in off-site offices 

 continue to manage EPCM contractor whose focus is now shifted mainly to field activity 

 initiate business readiness planning leveraging Owner’s Team resources, working collaboratively with the EPCM contractor 

 augment off-site human resources team 

 continue Mine Site development 

 continue delivery of P8 Quarry rock to WSD for construction 

  table continues…

 

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Year Activities
(Year -2)

 finalize concentrate smelting contract terms 

 continue to manage EPCM contractor whose focus is solely on field activity 

 establish satellite office and recruit port staff for facility at Stewart, BC 

 expand Terrace, BC office footprint to full size necessary to fully support mine and process operations on site 

 recruit all operations staff for HDS WTP to participate in wet commissioning of the facility following dry commissioning and handover by EPCM, initiate plant start up and operations 

 continue Mine Site development 

 continue delivery of P8 Quarry rock to WSD for construction 

(Year -1)

 recruit and on board all remaining positions for the process plant, metallurgical laboratory and TMF; plant operators are highest priority and TMF staff will be on boarded toward year end 

 complete recruitment for mine operations team 

 perform a significant amount of operator training in preparation for operations start up 

 oversight of OEM assembly of major mine fleet equipment 

 after completion of WSD and WTP, initiate ore mining in Mitchell open pit Phase I 

 delivery of first ore to Mitchell OPC 

 fully execute wet commissioning following dry commissioning and handover by EPCM of the process plant and all infrastructure 

 initiate ore transfer through the MTT controlled by process operations 

 introduce first ore to Treaty OPC and begin process plant ramp up 

Operations

 

Year 1

 

 complete recruitment and on boarding for remaining G&A and process operations staff during process plant ramp up 

 complete oversight of OEM assembly of major mine fleet equipment 

 produce first copper concentrate and doré at the Treaty OPC 

 initiate sand tailings production at the TMF and train employees on procedures for this long term dam construction effort 

 achieve commercial production 

 ramp up to full scale mining and primary crusher operations at the Mitchell OPC from Phase I mine development stages at Mitchell and Sulphurets 

 shipment of copper concentrate from the port facility in Stewart 

 

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19.0Market Studies and Contracts

 

Seabridge engaged NSA to provide opinion reports on marketing inputs for the 2016 PFS and review the 2019 copper-gold concentrate market and related concentrate treatment charges, excluding off-site transportation costs. In May 2022, NSA provides an update to this opinion focusing on the validity of the assumptions. The information and options in this section mainly come from the 2016 and 2019 Opinion Reports with applicable comments where appropriate.

 

No smelter contracts are currently in place or being negotiated. All currency amounts used in this section are in US dollars, unless otherwise specified.

 

19.1Copper Concentrate

 

19.1.1Marketability

 

When considering the marketability of copper concentrates, quality and quantity are determining factors. There is considerable variation in the quality of concentrates and the requirements of various smelters do vary; such variation relates to the technical abilities of the smelter and its overall concentrate feed and blend.

 

While smelters prefer a feed with about 30% copper and similar amounts of iron and sulphur, copper grades from many major high-grade suppliers have been falling and the market is seen the blend for many smelters dropping to a copper content of about 27%. Apart from the copper content, levels of iron, and sulphur, other key elements including gold and silver and impurity content are factors in concentrate market. Based on the impurity levels projected by Tetra Tech (using the test results completed to date – see Table 17.3), concentrates are relatively clean. Depending on the prevailing market at the time of contract negotiations, penalties will likely be minimal if any. Certain smelters in Japan, South Korea, and Europe, have more interest in copper concentrates with high gold content.

 

19.1.2Smelting Terms

 

Copper Concentrate Smelting Market

 

Copper concentrates account for approximately four-fifths of total newly-mined copper production, with the balance of output coming from solvent extraction and electrowinning copper cathode and other copper-bearing by-products.

 

Concentrate supply started to increase over 2013 and continued to increase to the current time with development of announced expansions of operating mines continuing. Integrated smelters process a significant portion of copper concentrate production. Such smelters are captive plants vertically integrated with mines through ownership. Non-integrated copper concentrate production globally is treated by custom smelters generally unintegrated with mines, although many may have investment ownership in mines. Custom smelters have increased their overall smelting market share following significant expansion of smelting and refining capacity, particularly in India and China. Such smelting industry has increased imports, as limited domestic mine capacity does not meet demand. This trend has been a key determinant in world concentrate supply/demand balances.

 

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The copper concentrate market has seen significant structural imbalances in the recent years between mine production and smelting capacities. Market treatment and refining charges have been volatile. Having fallen significantly to very low levels, at the time of writing there have been significant increases in smelter treatment charges and refining charges (TCs/RCs). Supply demand balance for concentrates determined by mine production relative to availability of capacity of the smelting industry. The availability of custom concentrates, relative to smelting capacity, should, in theory, be the ultimate determinant of terms for custom treatment of concentrates. However, the focus on strategic metals of which copper is one with the demand for EV’s has stimulated trade in copper concentrates and moved the copper price.

 

Copper Concentrates Contracts and Terms

 

The concentrate market consists of two types of contracts – longer term off-take contracts between mines and smelters generally reflecting the annual concentrate supply and demand balance and spot or short-term business primarily between mines and traders. Such spot business on a smaller scale exists between mines and smelters. By its nature, spot business is more volatile and there is considerable variation in spot TCs/RCs, not only annually, but over any year.

 

Current and Future Terms

 

NSA suggests that annual benchmark numbers are beginning to reflect a move towards sustainable long-term numbers. NSA believes that the most likely scenario is that ultimately charges need to move up towards a level that is economical for the smelting industry over the long term. Historical benchmark numbers are shown in Table 19.1

 

Table 19.1Benchmark Smelting Terms

 

  2017 2016 2015 2014 2013
Copper Treatment Charges ($/dmt) 92.50 97.50 107.00 92.00 70.00
Copper Refining Charges ($/lb) 0.0925 0.0975 0.1070 0.0920 0.0700
  2018 2019 2020 2021 2022
Copper Treatment Charges ($/dmt) 82.25 80.80 62.00 59.00 65.00
Copper Refining Charges ($/lb) 0.0823 0.0808 0.0620 0.0590 0.0650

 

In the 2016 Opinion, it was noted looking at the history of treatment refining charges over approximately 20 years up to 2005, these averaged about $77 per DMT and 7.7 cents per pound (this included approximately one cent of participation for these purposes split between the treatment charge and the copper refining charge) at an average price of $0.93 per pound of refined copper. With the copper price today at about four times as price, current treatment charges are well out of line with history. This could be said to support treatment charges needing to be higher in the future given cost increases.

 

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For comparison, the spot market April 2016 sales into the Chinese market the levels of TCs/RCs were between $90/dmt and $95/dmt of concentrate and $0.090/lb and $0.095/lb of copper, respectively. Today, spot treatment charges traders buying from mines are indicated as TC $70 to $80/dmt with RC $0.07 to $0.08/lb of refined copper. The comparative number for sales into China would be around TC $80 to $90/dmt with RC $0.08 to $0.09/lb of refined copper. The general view is to not expect price participation materializing near term, but it should not be ignored.

 

In 2016, a TC of $100/dmt coupled with any RC of $0.10/lb of refined copper was suggested. In 2019 and as a noted this was conservative when linked to copper price assumptions and for planning there was a case for reducing these assumptions to TC $95/dmt and RC $0.095 /lb of refined copper. Costs in both for mining and smelting are rising with par costs being particularly noteworthy. Inflation is putting pressure on labour costs and further upward movement is probable. The level of long-term TC/RC can be debated with a range being TC $90 to $95/dmt and RC $0.090 to 0.095/lb of refined copper in constant dollars.

 

TCs/RCs are not the only terms that are used in valuing copper concentrates. Payments and deductions are a matter of negotiation and will vary with many factors, including supply and demand, and custom individual markets.

 

The following terms are an indication of “standard” long-term smelter charges, including suggested TC/RC terms. Delivery is based on Cost, Insurance and Freight – Free Out (CIF-FO) smelter ports (the mine pays all costs up to delivery port and the buyer arranges and pays for cargo discharge).

 

Payable Metals

 

CopperPay 96.5% with a minimum deduction of 1 unit (amount deducted must  equate to a minimum of 1% of the agreed concentrate copper assay).

 

SilverIf over 30 g/dmt pay 90%.

 

GoldA scale is applicable with some variations of the following:

 

less than 1 g/dmt, no payment

 

1 to 3 g/dmt, pay 90%

 

3 to 5 g/dmt, pay 93%

 

5 to 7 g/dmt, pay 95%

 

7 to 10 g/dmt, pay 96.5%

 

10 to 20 g/dmt, pay 97%

 

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over 20 g/dmt pay 97.5%

 

over 30 g/dmt pay 97.75%.

 

Gold and silver payments may vary between smelter locations. In China, high gold in copper concentrates is not generally desired; relating more to internal pricing issues rather than technical concerns. Technically, the more modern smelting facilities are able to accept payment formulas similar to Japan and South Korea, but for many of the older smelters in North China, this is not the case. In Europe, with grades of over 40 g/dmt of gold content, payment of 97.75% with a minimum deduction of 1 g/dmt is likely to apply.

 

Refining Charges

 

Copper$0.095/lb payable copper

 

Gold$6.00 to $8.00/oz payable gold

 

Silver$0.50/oz payable silver

 

Treatment Charges

 

Treatment Charge       $95.00/dmt CIF-FO main smelter port.

 

Price Participation

 

Not applicable at present.

 

Penalties

 

Arsenic:$2.50 to $3.00 per 0.1% over 0.1% up to 0.5% arsenic

 

Antimony:$3.00 to $4.00 per 0.1% over 0.1% antimony

 

Lead:$2.00 to $3.00 per 1% over 0.5% to 1.0% lead

 

Zinc:$2.00 to $3.00 per 1% over 2% to 3% zinc

 

Mercury:$2.00 per each 10 ppm over 10 ppm mercury

 

Bismuth:$3.00 to $5.00 per 0.01% over 0.03 to 0.05% bismuth

 

Selenium:$3.00 to $5.00 per 0.01% over 0.05% selenium

 

Tellurium:$4.00 to $5.00 per 0.01% over 0.02% to 0.03% tellurium

 

Fluorine$1.00 to $2.00 per 100 ppm over 300 ppm fluorine

 

Chlorine$1.00 to $3.00 per 100 ppm over 300 ppm chlorine.

 

Furthermore, penalties may also vary from smelter to smelter. It should be noted that for the elements where a percentage range is used, this relates to ranges of penalty thresholds that are negotiated. The penalties noted in this section are generally in line with levels applicable over recent years, but there is a tendency towards higher levels.

 

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Based on the anticipated impurity levels derived from the test results by Tetra Tech (as presented in Table 17.3), the concentrates from the KSM concentrate production are relatively clean, and depending on the market situation at the time of contract negotiations, penalties will likely be minimal if at all applicable. As most of the mill feeds will be the blended materials from different deposits and spatial locations, the blend should effectively mitigate penalty elements rising for the ore from some limited locations.

 

Other Off-site Costs

 

Various indirect costs other than smelter charges include:

 

Losses; assumed to be 0.1% or less, due to improvements in material handling.

 

Insurance; marine insurance is assumed to be in the range of 0.1 to 0.15% of net invoice value of the concentrate.

 

Supervision, assaying and umpire costs; the costs for third-party supervision and assaying are assumed to be approximately US$1/dmt.

 

Marketing; the cost of marketing varies with concentrate tonnage, location, and number of smelters to be shipped. For 2022 PFS, the estimated marketing cost is in the range of US$5 to US$10/dmt.

 

Concentrate transportation; the transportation costs for copper concentrate are based on the following assumptions by Tetra Tech:

 

qtrucking: US$35/wmt

 

qport storage and handling: US$17/wmt

 

qocean transport to Asian port: US$49/wmt.

 

19.2Molybdenite Concentrate

 

19.2.1Smelting Charge

 

Molybdenum concentrates of either primary production origin, or as a co-product, need to be further processed. This is initially to produce molybdenum oxide by roasting, or by use of autoclaves for upgrading. Quality is an important consideration and certain elements can be deleterious. As a rule of thumb, 50% molybdenum content is considered the minimum. Below that, buyers will begin to be a bit selective and charges will rise somewhat. One guideline, given for each 1% below 50%, would be an increase in charges of $0.05/lb of molybdenum.

 

Currently, the deduction for roasting is very quality dependent, with high copper content concentrates generally selling at a 10 to 15% discount. Assuming that the copper content of its molybdenum concentrates is reduced to 0.45% or less, the lower end of the range will apply to clean high-grade concentrates. In the years (2005 to 2009), discounts for high-copper molybdenum concentrates have reached 25%.

 

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In summary, it is recommended to use a discount of 12% from the price, with a minimum of $1.00/lb and a maximum of $2.50/lb. This discount would be inclusive of all charges mine to market, such as delivery costs, irrespective of whether or not the concentrate is sold to a trader or a roaster directly.

 

On average, the molybdenum concentrate from KSM could contain approximately 1,000 to 2,000 ppm rhenium or higher. Given the high rhenium content, some roasters would recognize the rhenium content in the form of lower treatment charge or some payments. However, it is difficult to project the rhenium value at the current market and the level of study. This potential for additional value in the rhenium content represents a future opportunity.

 

19.3Gold and Silver Doré

 

There are no gold and silver doré marketing studies conducted for this PFS. No smelter contracts are currently in place or being negotiated. The terms used for this study are based on the typical terms currently used in the markets. All currency amounts used in this section are in US dollars, unless otherwise specified. The general payment terms for gold and silver doré are assumed as below:

 

Gold: pay 99.8% of content less a refining charge of US$1.00/accountable oz. Doré transportation is assumed to be US$1.00/oz.

 

Silver: pay 90.0% of content less a refining charge of US$1.00/accountable oz. Doré transportation is assumed to be US$ 1.00/oz.

 

Insurance and assay costing is assumed to be in the range of 0.10 to 0.15% of net invoice value of the doré.

 

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20.0Environmental Studies, Permitting, and Social or Community Impact

 

20.1Licensing and Permitting

 

The KSM mine development plan was subject to the BCEAA, the Canadian Environmental Assessment Act- 1992 (CEAA), and Chapter 10 of the NFA. In this section of the Report, the term “Project” refers to a mine development plan that would substantially support what is contemplated in the 2022 PFS.

 

As of May 2022, the Project has completed the provincial and federal environmental assessment review processes, and the appropriate certificates/approvals have been obtained for this stage of the Project’s development. Additionally, permits for early-stage construction activities, continuation of exploration, certain permit and project approval renewals have also been obtained from provincial agencies, and authorizations to construct and operate obtained from the federal agencies. KSMCo continues to advance permitting to allow for the construction of the KSM mine, as well as to continue exploration activities. Details of the provincial, federal, and NFA processes and current status, as well as the current permitting status of KSM property, are included in this section.

 

The KSM Project completed a harmonized EA process with the provincial and federal governments, in accordance with the principles of the Canada-BC Agreement on Environmental Assessment Cooperation (Cooperation Agreement 2004). The KSM Project also completed a separate EA review and received approval as required under the Nisga’a Final Agreement. The harmonized EA process included a working group comprising federal and provincial officials, the Nisga’a Lisims Government (NLG), Indigenous groups, and local government agencies. Although the KSM project does not require authorizations from US agencies, the Project EA involved representatives of the US federal and Alaska state agencies due to the project’s location near the US/Canada border.

 

The following major permits have been obtained for the KSM property:

 

BC EMLI Mines Act (1996) Permit M-245 (amended 17 May, 2022 - amalgamating Mines Act Permits MX-1-571 (KSM), MX-1-763 (PTMA), MX-1-965 (DKEA), MX-1-998 (KSM Iron Cap High Camp), MX-100000044 (Snowfields).

 

Mining Leases 1031440 and 1031441 were issued to Seabridge Gold Inc. on October 6, 2014.

 

BC Ministry of Forests, Lands, Natural Resource Operations and Rural Development (BC MFLNRORD) Land Act (RSBC 1996c), Licence of Occupation for MTTs Roadway SK904033. Licence of Occupation for Treaty Transmission Line SK908555.

 

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BC MFLNRORD, Forest Practices Code of British Columbia Act Special Use Permits for Treaty Creek (S25751) and Coulter Creek Access Roads (S25750).

 

BC MFLNRORD, Forest Act Occupant Licence to Cut L49546 (Mine Site), L49608 (CCAR), L49658 (PTMA) and L49612 (TCAR).

 

BC ECCS, Environmental Management Act, Authorization for Effluent Discharge PE106814 (TWTP#6), PE108155 (DKEA), and PE106824 (TWTP#4).

 

BC ENV, Environmental Management Act, Authorization PA 106826 for air emissions for five KSM camps.

 

BC ENV, Environmental Management Act, Authorizations PR 106834 (Mine Site) and PR106835 (PTMA) for authorized landfills.

 

BC ENV, Environmental Management Act, Registration for Municipal Wastewater Regulation Authorization PE 106836 for waste discharge to the environment for Camp 9/10 and PE106837 (Camp 6), PE106839 (Camp 5), EP106809 (Mitchell Operating Camp), PE106841 (Camp 4).

 

Canada Department of Fisheries and Oceans, Environment Canada, Fisheries Act Authorization under Paragraphs 34.4(2)(b) and 35(2)(b) of the Fisheries Act for mine waste discharge into fish habitat resulting in a harmful alteration, disruption, or destruction (HADD), and the authorization of fish habitat offsetting works (Glacier, Taft, and Treaty creeks).

 

Transport Canada, Canadian Navigable Waters Act, Authorization to construct a bridge over the Bell-Irving River for the KSM Mine Access Road.

 

Environment Canada, International River Improvements Act, License to construct and operate improvement works on tributaries of the Unuk River which is an international river crossing the Canada-US border.

 

Nisga’a Final Agreement (NFA), Chapter 10 – Environmental Assessment and Protection approval in December 2014

 

20.1.1Provincial EA Process

 

Under the BC EA process, certain projects must undergo an EA, and an EA Certificate must be obtained before the KSM mine development can proceed. The scope, procedures, and methods used for each assessment are tailored to the specific circumstances of a proposed project. The EA must assess a project’s potential environmental, economic, cultural, social, heritage, and health effects.

 

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The KSM Project officially began the BC EA process in March 2008, with the submission of a Project Description (Rescan 2013). Over the following six years, the project underwent parallel provincial and federal EA processes. In July 2013, KSMCo submitted an Application/EIS (Rescan 2013) under the BC Environmental Assessment Act (BCEAA 2002) to the BC Environmental Assessment Office (BCEAO). EA Certificate #M14-01 for the KSM project was issued on July 29, 2014 (Certificate). The Certificate required that the KSM Project be substantially started within five years of July 29, 2014.

 

The full Application/EIS can be found on the BCEAO web site: http://a100.gov.bc.ca/appsdata/epic/html/deploy/epic_project_doc_list_322_r_app.html

 

On March 21, 2019, the date specified in the Certificate by which KSMCo must have substantially started the Project was extended from July 29, 2019 to July 29, 2024.

 

In August 2020, KSMCo applied for an additional two-year extension of the Certificate from July 29, 2024 to July 29, 2026 due to the impacts of the COVID-19 pandemic. On November 16, 2021, the Minister extended the date specified in the Certificate by which KSMCo must have substantially started the Project, to July 29, 2026.

 

20.1.2Federal Process

 

The KSM Project was subject to a comprehensive study level of assessment under the CEAA (1992).

 

A “comprehensive study” of the KSM Project commenced in July 2009, and the federal EA process continued over the following five years along with the provincial EA process. The KSM (Kerr-Sulphurets-Mitchell) Project Comprehensive Study Report (KSM Project Comprehensive Study Report) was issued by the Canadian Environmental Assessment Agency in July 2014.

 

The KSM Project Comprehensive Study Report, along with comments from the NLG, other Indigenous groups, and the public, were considered by the Minister of the Environment when making her final EA decision. KSM received federal approval that the Project could proceed on December 19, 2014.

 

Nisga’a Final Agreement

 

The NFA is a treaty signed by Nisga’a Nation, the Government of Canada, and the Government of BC in 1999.

 

KSMCo proposes to develop some components of the 2022 PFS footprint within the Nass Area, including the Treaty OPC, the TMF, and the northern portion of the MTT. No components of the KSM infrastructure will physically occupy any portion of Nisga’a Lands or the Nass Wildlife Area (NWA), both of which are located south of the potentially affected portion of the Nass Area.

 

The NFA provides for Nisga’a participation in federal or provincial EAs of projects within the outer Nass Area boundary and an assessment of potential adverse project effects on residents of Nisga’a Lands, the Nisga’a Lands themselves, or more generally, on Nisga’a interests as set out in the NFA.

 

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The Government of Canada worked with the NLG and the Government of BC to facilitate the assessment of Nisga’a effects as part of the CEAA comprehensive study. KSMCo completed an economic, social, and cultural impact assessment (ESCIA) on the well-being of Nisga’a citizens based on a work plan that was required by the joint Application Information Requirements (AIR). Effects were described in the Application/EIS as part of KSMCo’s analysis of the effects KSM mine development will have on environmental valued components (VCs).

 

The KSM Project Comprehensive Study Report examined effects on Nisga’a citizens, lands, and interests. KSM received its NFA approval in December 2014 following receipt of the Federal CEAA Approval on December 19, 2014.

 

20.1.3Provincial Permits

 

The Application/EIS was accompanied by applications for eligible provincial authorizations in accordance with the BC Environmental Assessment Act (BCEAA 2002) Concurrent Approvals Regulation (BC Reg. 371/2002).

 

This set of initial permits is referred to as the “Batch 1 Permits” and included permits for the following mine components:

 

KSM Project Mines Act and Environmental Management Act Permit Application for Limited Site Construction (May 2013)

 

Special Use permits for the CCAR and TCAR

 

KSM Construction Camps

 

KSM Project Treaty Transmission Line

 

Land Act tenure for the MTT Roadway.

 

In November 2015, KSMCo submitted a Mines Act Notice of Work and an Environmental Management Act (2003) permit application to construct the Kerr Exploration Adit (DKEA). Mines Act permit MX-1-965 authorizing construction of DKEA was issued September 20, 2016, and EMA permit PE-108155 was issued on August 24, 2016 authorizing the discharge of mine effluent from the DKEA to Sulphurets Creek. The DKEA will be located near the existing temporary KSM exploration camp.

 

As of May 2022, KSMCo estimates that upwards of 75 provincial permits, licences, and approvals may be required to fully develop the 2022 PFS. As KSMCo undertakes further project optimization studies, permit requirements are routinely evaluated for the need to amend existing issued permits, or to identify where new permits, licences, and approvals may be required to take into account project optimization initiatives. A single consolidated Mines Act Permit is being developed for the Project to make more efficient permit condition management. A register of permits is maintained.

 

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20.1.4Federal Permits

 

The Application/EIS included applications for the following federal permits and authorizations:

 

Metal and Diamond Mining Effluent Regulations (MDMER) Schedule 2 Amendment

 

International Rivers Improvements Act Licence

 

Canadian Navigable Waters Act authorization.

 

Further details are provided in the following sections.

 

Fisheries Act (RSC 1985, c F-14) - MDMER Schedule 2 Amendment Application

 

Section 36(3) of the Fisheries Act (1985) prohibits the deposit of deleterious substances into water frequented by fish (i.e., fish-bearing waterbodies); however, deposition of metal mine tailings and/or waste rock into such waterbodies may be exempted from this prohibition if the water bodies are included in Schedule 2 of the MDMER of the Fisheries Act (1985) (Schedule 2). In order to permit the construction of the proposed tailings management facility in North Treaty Creek and South Teigen Creek (both fish-bearing streams), an application to amend Schedule 2 of the MDMER was submitted with the Application/EIS along with a fish habitat compensation plan designed to offset the habitat losses associated with the Schedule 2 amendment (see MMER Compensation Plan – Appendix 15Q of the Application/EIS). The waterbodies contained in the proposed KSM tailings impoundment area were gazetted and added to Schedule 2 in July 2017.

 

Other Fisheries Act Permitting

 

Section 34.4 of the Fisheries Act prohibits any work that results in the death of fish, unless authorized. Section 35 of the Fisheries Act prohibits the HADD of fish habitat unless authorized. KSM requires an authorization under Sections 34.4 and 35(2) of the Fisheries Act to permit the HADD associated with other Project infrastructure.

 

On August 4, 2021, Fisheries and Oceans Canada issued an authorization under Paragraphs 34.4(2)(b) and 35(2)(b) of the Fisheries Act to KSMCo for the proposed project works, undertakings and activities associated with the TMF construction and associated diversions, and installation of Treaty Creek pipeline outlet. The term of the authorization extends to December 31, 2029, and may be renewed. The Fisheries Act authorization includes Condition 4.4 requiring the construction of the Glacier Creek Fish Habitat Offsetting Plan (FHOP). Construction of the Glacier Creek FHOP was initiated in 2021 and is projected to be completed in 2022. Additional fish habitat offsetting plans required under Schedule 2 are planned to begin construction for Taft Creek in 2022/2023, and for Treaty Creek in 2024.

 

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International River Improvements Act licence

 

An application for a licence under the International River Improvements Act was submitted to Environment Canada in February 2015.

 

The application was prepared in accordance with Sections 6 and 7 of the International River Improvements Regulations (C.R.C., c982). The application is for improvements on the Unuk River, specifically within the Sulphurets Creek Watershed (which is a tributary to the Unuk River) (Improvements).

 

The licence was issued on October 21, 2016, and transferred to KSMCo in 2018. The licenced Improvements include the Water Storage Facility and ancillary water works for 2022 PFS, representing dams, reservoirs, and associated water diversion, collection, and management structures for the purpose of: diverting fresh (non-contact) water around the mine site to downstream receiving waters and collecting water that has been in contact with disturbed areas from the mine site for control prior to discharge into the receiving waters.

 

The Licence is valid for 25 years until October 20, 2041.

 

Navigable Waters Approvals

 

Further to a letter received from Transport Canada dated August 1, 2014 (where Transport Canada determined that waterways that were assessed as part of the EA process were not navigable waters), Transport Canada subsequently determined that the Bell Irving River is navigable for the purposes of the Canadian Navigable Waters Act, and an approval was required for the construction of the Bell Irving River Bridge. KSMCo submitted the application on July 12, 2021, and received approval from Transport Canada on November 18, 2021, which authorized a Permanent Bridge and Temporary Bridge crossing the Bell Irving River. An approval will also be required for the Unuk River crossing for the Coulter Creek Access Road.

 

20.1.5benefits agreements

 

Nisga’a Nation

 

On June 16, 2014, KSMCo entered into a comprehensive Benefits Agreement with the Nisga’a Nation in respect of the Project. The Benefits Agreement establishes a long-term co-operative working relationship between KSMCo and the Nisga’a Nation, under which the Nisga’a Nation will support the mine development and participate in economic benefits from the Project Environmental Settings and Studies.

 

The Project is located in the coastal mountains of northwestern BC, approximately 950 km northwest of Vancouver, 65 km northwest of Stewart, and 35 km northeast of the BC-Alaska border.

 

The following sections summarize the valued components of the biophysical and socio-community aspects of the Project that have been studied and reported in the Application/EIS (Rescan 2013) in relation to the EA process. Studies occurred at different scales involving detailed studies within a Local Study Area (LSA) and lower density studies in a Regional Study Area (RSA) as defined in the Application/EIS.

 

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Tahltan Nation

 

On July 8, 2019, the Tahltan Nation and KSMCo announced that the parties reached agreement on the terms of a Co-operation and Benefits Agreement in connection with the Project (IBA). The IBA provides a framework for the parties to work together in relation to the Project and includes commitments to economic benefits and environmental management of the land.

 

20.1.6Biophysical Setting

 

Air Quality

 

The air quality in the area proposed for the Project is predominantly unaffected by anthropogenic sources, reflecting the region’s remoteness and the lack of, and localized nature of, sources of anthropogenic air emissions.

 

Geology and Geochemistry

 

The mineralized zones in the local area and more regionally, tend to be sulphide-rich. Where sulphide minerals such as pyrite are present, oxidation can create ARD, unless sufficient quantities of neutralizing minerals are available. In the event that acidic drainage is formed, low pH conditions can lead to higher rates of metal leaching (ML). Baseline surface water and groundwater quality in the vicinity of mineralized zones in the region exhibit relatively low pH and significant metal concentrations, reflecting the presence of sulphide minerals and the natural occurrence of ML/ARD processes.

 

Physiography

 

The KSM topography is very rugged. Glaciers are common in high elevations. Most steep slopes consist of exposed bedrock and accumulations of rubbly colluvium. Gentler slopes have a thin mantle of morainal material (glacial till). Thick glacial deposits are generally restricted to the margins of major valley floors and adjacent lower slopes. Avalanches and slope failures are common features at high and intermediate elevations (above 1,500 masl).

 

Geohazards

 

Locally and regionally, geohazards are linked primarily to landslides and snow avalanches. Landslide hazards are abundant throughout the region. They are attributed to several factors, including the presence of unstable surficial soils and weak bedrock, repeated geologically recent glaciations, resulting in over-steepened valley sidewalls, the loss of slope buttress support following glacial recession, abundance of veneers that are shallow to bedrock, and the high precipitation environment.

 

Unstable lateral morainal till has been deposited on slopes at angles that exceed the angle of repose, resulting in rubbly colluvium accumulating along moderate steep slopes and valley bottoms. The unloading of the valley walls following glacial retreat has led to pressure release cracks and associated local instability on over-steepened slopes resulting in geohazards, such as rock fall, debris avalanches, and slumping of surficial materials. These geohazard processes are endemic to the local area.

 

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Snow avalanche hazards are abundant due to high elevation, substantial snow supply and generally steeper slope gradients, and tend to be associated with terrain that is open and steep.

 

Hydrology/Surface Water Quantity

 

Regional and local surface water quantity characteristics were determined from data collected from specially installed hydrometric stations, used in conjunction with a regional analysis prepared for long-term hydrometric data from Water Survey of Canada hydrometric stations.

 

The monthly distribution of flow tends to be concentrated in the open water season (May to October), with less than 20% of the annual flow occurring from November to April at a majority of the regional stations.

 

Groundwater Quantity

 

Groundwater conditions correspond with the mountainous, wet environment that comprises the mine site and the PTMA. Groundwater gradients are high, driven by heavy rainfall and recharge at higher elevations in the mountains. Valley bottoms are discharge zones, with groundwater levels near or above (artesian) ground surface. Discharge zones also exist along valley walls in the mine site, where seeps of acidic water have been observed (with pH readings as low as 2.5).

 

Surface Water Quality

 

The hydrological regime affects water quality in two ways:

 

increased flows during freshet, glacial melt, and heavy rainfall events dilutes concentrations of major ions and total dissolved solids

 

increased sediment load and transport during high-flow periods leads to increased concentrations of TSS and particle-associated metals.

 

Streams near the mine site and PTMA have distinct surface water quality. Metal leaching due to naturally occurring ARD is associated with total and dissolved metal concentrations in Mitchell and Sulphurets creeks that are frequently higher than levels established in BC water quality guidelines for the protection of freshwater aquatic life. The high suspended sediment load, low concentrations of bioavailable nutrients and high concentrations of total and dissolved metals identified in Mitchell and Sulphurets creeks are contributing factors to the poor productive capacity of mine site streams. The lower suspended sediment load, increased concentrations of bioavailable nutrients, and lower concentrations of total and dissolved metals identified in the Snowbank, Teigen, Treaty, and Bell-Irving watersheds are contributing factors to the greater productive capacity of PTMA streams relative to the mine site.

 

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Groundwater Quality

 

Groundwater quality at the mine site is heavily influenced by the sulphide ore deposits. Groundwater is acidic near, and within, the mineral deposits, with pH measurements as low as 2.5 in seeps along the valley walls of Mitchell Creek. Concentrations of certain metals are elevated in groundwater throughout the mine site, and are particularly high near and within the mineral deposits. Metals with elevated concentrations include iron, aluminum, copper, chromium, lead, manganese, and zinc. Groundwater in the Mitchell Valley is not suitable for human consumption or the sustenance of fresh water aquatic life.

 

Wildlife Species

 

Mature forests, wetlands, alpine areas, and riparian forests provide high-value habitat to a diverse wildlife community. Common species or groups that occur in the RSA include ungulates (e.g., moose and mountain goat), omnivores/carnivores (e.g., grizzly bear, black bear, and wolves), furbearers (e.g., fisher, marten, and wolverine), hoary marmots, bats, birds (forest birds, raptors, and waterfowl), and amphibians (e.g., Columbia spotted frog and western toad). Forest harvesting within the RSA has been minimal compared to many other areas in BC, due to the remoteness of the area and the relatively poor productivity of the forests, so that the wildlife habitats found in the majority of the wildlife RSA are essentially undisturbed.

 

20.1.7Economic, Social, and Cultural Setting

 

Governance

 

There are five levels of governance in the area of northwestern BC where the 2022 PFS will be developed. Municipal, regional, provincial, and federal bodies comprise the non-Indigenous forms of governance, while Indigenous communities have their own governing bodies.

 

The 2022 PFS is situated in the Regional District of Kitimat-Stikine, and Electoral Area A of the Bulkley Nechako Regional District. Local communities include municipalities, Nisga’a villages, other Indigenous communities, and unincorporated settlements. Municipal governance only exists for the District of Stewart, the City of Terrace, the Village of Hazelton, the District of New Hazelton, and the Town of Smithers. The remaining communities that are not administered by Indigenous bodies (Dease Lake, South Hazelton, Bell 2, Meziadin Junction, and Bob Quinn Lake) are unincorporated and governed by the regional district in which they are situated.

 

Economic Setting

 

Economically, the region where the Property is located has been dependent upon timber and minerals for well over 100 years. The majority of non-Indigenous communities in the region were initially established to serve natural resource activities such as the mine operations near Cassiar, Stewart, Smithers, and Bob Quinn Lake. Historically, the region’s economic and social diversity has been constrained by limited access and infrastructure, lengthy distances, remote and small communities which provide limited labour or services, and long winters. More recently government and private investment has improved road access, port development, and independent power generation, and brought provincial grid power into some parts of the region. The Red Chris and Brucejack mines recently opened, providing long-term direct and indirect employment to local communities and others within and outside of the region.

 

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Forestry, fishing, and mining were the key economic drivers of northwestern BC through the 1950s to the 1980s. Today, the economies of local communities continue to be largely resource-based, and focus on supporting these sectors in the region.

 

Overall, the economy in northwestern BC is gradually becoming more diversified and now includes hydroelectric power generation. In some communities, employment levels have increased in the public service, sales and service, tourism, transportation, and mineral exploration sectors. Employment sectors in local Indigenous communities now include sales and service, mineral exploration, labour, and government administration components. Expanded education and training opportunities to support regional economies is also delivered regionally, enabling residents to train and work locally. There are recent signs that the population decline may be reversing.

 

Today, the mining industry continues to provide an important source of employment in the region, supplying an estimated 30% of jobs for communities along Highway 37 in recent years.

 

Social Setting

 

Recent economic conditions in the mining sector influence population levels in regional communities. Population losses have stabilized and are now increasing with the opening of Brucejack and Red Chris mines, hydroelectric power operations, port development and a healthy mineral exploration industry.

 

Indigenous Groups

 

Several Indigenous groups may be potentially affected by the development of KSM. The PTMA is situated within the Nass Area, as defined by the NFA. The Tahltan Nation (as represented by the Tahltan Central Council) asserts a claim over part of the 2016 PFS footprint. The Gitanyow First Nation (notably wilp Wiiltsx-Txawokw), the Gitxsan Nation (as identified by the Gitxsan Hereditary Chiefs Office), and wilp Skii km Lax Ha have identified potentially affected interests within the broader region, notably downstream of the PTMA. The Skii km Lax Ha are also asserting claim of an area covering the mine site and PTMA.

 

Land Use Setting

 

The Property is located in an area of northwestern BC known as the “Golden Triangle”, due to its high mineral potential and the occurrence of several gold projects in the region. For the past century, land and resource uses in the region have been largely driven by forestry, mining, and mineral exploration, and this is still true today. A limited amount of commercial and non-commercial recreation also occurs in the region, including hunting, trapping, fishing, heli-skiing, hiking, and camping.

 

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Land Use Planning Context

 

The Project area is subject to the provisions of two land use plans—the Cassiar Iskut-Stikine Land Resource Management Plan (LRMP) and the Nass South Sustainable Resource Management Plan (SRMP), developed in partnership with Indigenous groups, government, and non-government interests.

 

Mineral resource activity, timber harvesting, commercial recreation and tourism, guide outfitting, hunting, fishing, trapping, and cultural land uses are all allowable activities within these land use plans.

 

20.2Water Management

 

20.2.1Overview of Water Management

 

An extensive system of water management facilities will be constructed and maintained throughout the life of the KSM mine to divert fresh (non-contact) water away from disturbed areas and to collect water that has contacted disturbed areas (contact water) for treatment before release into the environment. Please refer to Section 18.2 of this Report for details of the updated water management plans for KSM.

 

An overview of the Water Management Plan for operations is included in Figure 20.1.

 

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Figure 20.1KSM Mine Site Water Management Schematic

 

 

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20.2.2Summary of Water Management Plan

 

Objectives and Targets

 

The objectives of the Water Management Plan are to provide a basis for management of surface water on site including:

 

diverting non-contact water around the mine site and PTMA

 

protecting ecologically sensitive areas and resources and avoiding harmful impacts to aquatic life and wildlife habitat

 

providing and retaining water for mine operation

 

defining required environmental control structures

 

·collecting and treating contact water from the mine site to meet discharge requirements prior to release to the receiving environment.

 

The targets intended to optimize the Water Management Plan to achieve surface water objectives include:

 

minimizing the production of contact water by implementation of best management practices and water diversion measures

 

collecting and treating contact water where required in order to meet applicable water quality standards

 

·implementing and maintaining an on-site monitoring and control system to regulate surface water quantity and quality.

 

Legislation and Standards

 

The Water Management Plan has been developed in accordance with the following legislation:

 

BC Mines Act (1996)

 

Fisheries Act (RSC 1985, c F-14)

 

Canada Water Act (1985, c. C-11)

 

BC Environmental Management Act (2003)

 

BC Water Sustainability Act (2014)

 

·BC Water Protection Act (1996)

 

Processing and Tailing Management Area

 

Water management in the PTMA is focused on the construction of diversion channels to control and divert water in the PTMA catchment area to either South Teigen Creek or North Treaty Creek.

 

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Monitoring

 

Monitoring programs will enable KSM mine operators to measure the success of its management strategies and to identify where additional mitigation is necessary.

 

Several management plans and monitoring programs include components that will assist in ensuring the long-term protection of the aquatic environment downstream of Property. Management plans and monitoring programs will be reviewed and updated as required.

 

20.3Waste Management

 

20.3.1Tailing Management Facility Management and Monitoring Plan

 

Tailing produced after the mined ore has passed through a process of high pressure grinding, flotation, and leaching at the Treaty OPC will be transported by slurry pipeline to the TMF. Conventional flotation tailing will be stored in two cells, the North Cell and the South Cell. A separate tailing stream consisting of sulphide-rich CIL residue tailing, produced as a waste product in the cyanide leaching process, will be transported in a pipeline to a fully lined CIL Lined Pond located in the Centre Cell between the North and South cells, and will operate during the filling of the North and South cells. The TMF is designed and approved under the environmental assessment approvals to store 2.29 Bt of tailing produced over the 35-year mine life.

 

The TMF will ultimately consist of three storage cells retained by four compacted cyclone tailing dams: the North Dam, the Splitter Dam, the Saddle Dam, and the Southeast Dam. Seepage from the tailing dams will be collected in seepage collection ponds constructed downstream of the tailing dams.

 

The tailing dams and associated seepage recovery dams proposed for the 2022 PFS fall into the category of major dams. Prior to commencing work, plans, operating, and monitoring procedures, developed in concert with an Independent Geotechnical Review Board, must be submitted for approval by the Chief Inspector of Mines. The Code also requires supporting plans to address ML ARD, Reclamation and Closure, Emergency Preparedness and Response, and closure cost estimates. Please refer to Table 22.5 for the cost.

 

Monitoring

 

A monitoring program will be developed that will include requirements for inspection of dams and water control structures and procedures for instrumentation monitoring during the construction, operation, closure, and post-closure phases. The results of the monitoring program will be reviewed on a regular basis to measure the success of the management strategies, to compare the recorded data against design criteria, and to identify where design changes or additional mitigation may be necessary.

 

A schedule for routine inspection and instrumentation monitoring will be developed at the time of mine permitting based on the mine construction and operation schedule. Additional inspections of the dams and water control structures will be undertaken following extreme rainfall events, significant runoff events, or significant earthquake events.

 

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The TMF will remain in operation after the end of mining operations until such time as the quality of the water stored in the impoundment reaches an acceptable level for discharge and reclamation is completed. The dam and associated facilities will require ongoing monitoring and maintenance to ensure dam safety and to meet regulatory requirements for dam safety.

 

In the event of temporary mine closure, visual inspection and maintenance of the dams, diversion channels, collection ditches, and spillways will be required.

 

20.3.2Best Available Tailings Technology Assessment

 

KSMCo recently completed an assessment of tailing technologies, tailing facility locations, and management practices. The assessment was an update of the tailing alternative assessment that was completed as part of the Application/EIS and the Schedule 2 Amendment process and subsequently re-completed in 2015–16 to ensure that the appropriate tailing technology had been selected (see Chapter 33 and Appendix 33-B of the Application/EIS; https://ksmproject.com/bat-report/ and KCB 2016a).

 

20.3.3Waste Rock Management

 

The proposed management of waste rock is outlined in the Rock Storage Facilities Management and Monitoring Plan, which can be found in Volume 26 of the Application/EIS (Rescan 2013).

 

Waste rock from open pit and underground mining operations that is not used for construction purposes will be consigned to the Mitchell RSF located in the Mitchell Creek Valley. All waste rock placed in the Mitchell RSF is assumed to be potentially acid generating. The total amount of mine waste rock to be removed during open pit excavations at the Mitchell/East Mitchell pit, and Sulphurets pit, is approximately 3 Bt over the LOM.

 

NPAG mine waste rock removed from the Mitchell P8 Quarry on the north Mitchell pit wall during pre-production will be used to construct the basal drain beneath the Mitchell RSF, and it will be used as rockfill material in the construction of the WSD.

 

The RSF proposed for 2022 PFS fall into the category of a major dump as defined under Section 10.5.5 of the Code (BC EMLI 2021).

 

A set of detailed mine development and reclamation plans will be submitted at a later date as part of the Mine Plan and Reclamation Program Permit application. The following sections outline the general provisions included in these documents in terms of RSF operation and monitoring.

 

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Closure

 

For long-term stability and reclamation purposes, the slopes of the Mitchell RSF will be re-contoured at closure. The modifications to the RSF areas will include:

 

re-contouring the Mitchell slopes below an elevation of 1,100 m (treeline) to an overall slope of 2H: 1V (27°) with a 50 cm soil cover and vegetated

 

re-contouring the western face of the Mitchell RSF below an elevation of 840 m

 

above the treeline, higher benches are proposed to be left un-vegetated to reflect existing talus slopes present in the Mitchell valley.

 

Please refer to the Application/EIS (Volume 26) (Rescan 2013) for further details.

 

Post-closure

 

The post-closure phase includes complete reclamation of the RSF and continued treatment of water collected in the WSF until the water quality meets acceptable standards for direct discharge to the environment. At that time, all facilities will be decommissioned and flows downstream of the mine site will be restored to pre-mine conditions.

 

20.3.4Domestic and Industrial non-hazardous and hazardous Waste Management

 

The proposed management of domestic and industrial waste, including hazardous and non-hazardous waste and dangerous goods, is outlined in the Domestic and Industrial Waste Management Plan and Dangerous Goods and Hazardous Materials Management Plan, (Waste Management Plans) which can be found in the Application/EIS (Volume 26) (Rescan 2013).

 

Landfills

 

Two landfills will be established, one in the Mitchell OPC and one in Treaty OPC. The landfills will be used to dispose of only solid inert, non-reactive waste such as used conveyor belts, empty dry latex paint cans, grinding balls, air filters, non-recyclable plastics, and incinerator ash. To deter wildlife attraction to the landfill, the landfill will be fenced and only solid inert waste that will not act as a wildlife attractant will be deposited there. The waste will be periodically covered with not potentially acid generating waste rock or local till to prevent wind loss and to mitigate wildlife attraction.

 

Hazardous Waste

 

Hazardous waste will be produced in all phases. It includes materials such as waste oil, laboratory chemicals and solvents, lead-acid batteries, oil filters, and used oily rags and absorbent pads.

 

The Hazardous Waste Regulation (BC Reg. 63/88) under the Environmental Management Act (2003) defines “hazardous waste” (see Volume 26 of the Application/EIS for additional details). Any updates on hazardous waste management since the Application/EIS can be found in Section 18.0 of this Report.

 

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Closure and Decommissioning

 

Activities during the closure phase will be similar to the activities during the construction phase. A range of materials will become available for salvage, recycling, or disposal with the dismantling and removal of buildings, surface structures, fuel tanks, etc. The Reclamation and Closure Plan will cover the closure, reclamation, and decommissioning of the mine infrastructures in detail.

 

Waste Management during Closure

 

Significant amounts of waste will be generated from the dismantling of buildings and process-related materials. The approach for waste management during closure will be to identify feasible salvage and recycling options.

 

Upon closure, the buildings, facilities, and process equipment will be dismantled and either disposed of at the site landfill (inert non-reactive materials only) or removed from the site for recycling and/or disposal. Any equipment or materials with market value will be removed for capital recovery.

 

20.4Air Quality Management including Greenhouse Gases

 

The proposed management of air quality, including greenhouse gases, is outlined in the Air Quality Management Plan and Greenhouse Gas Management Plan, which can be found in the Application/EIS (Volume 26) (Rescan 2013). Under the Canadian Environmental Protection Act (CEPA) the operating mine will be required to report to the federal National Pollutant Release Inventory program.

 

20.5Environmental Management System

 

KSMCo has developed a conceptual Environmental Management System (EMS) and associated Environmental Management Plans for KSM. As stated in the KSM Project AIR document approved by the British Columbia Environmental Assessment Office (BC EAO; 2011), Environmental Management Plans are essential for the EMS for any major development.

 

An EMS is a requirement of a Mines Act Permit for mines in BC and is the high level framework supporting each Environmental Management Plan. Environmental Management Plans are the specific and detailed goals, objectives, and procedures for the protection of worker health and safety, environmental monitoring, and operating procedures that show the regulatory agencies how legislation and regulations will be met at the mine site and the PTMA. Environmental Management Plans are managed collectively under the umbrella of the EMS.

 

Environmental Management Plans are to be applied during the planning, construction, operation, closure, and post-closure phases of KSM.

 

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20.6Closure and Reclamation

 

Closure and reclamation planning for KSM will contribute to the success of closure and reclamation during mining and at the end of mine life, which will reduce the need to restructure the infrastructure components, limit the amount of material re-handling, and reduce the environmental effects. Mine development and operation will incorporate techniques to minimize surficial disturbance and, where possible, progressively reclaim areas affected during construction and operation. Stabilizing and rehabilitating surfaces will reduce the potential for degradation of the resources due to extended exposure to climatic factors, reducing closure-related capital costs at the cessation of mining activities. The final mine closure plan will be developed during mine operations prior to closure with full involvement of regulatory agencies, Indigenous communities, and other stakeholders.

 

20.6.1Closure and Reclamation Objectives

 

The conceptual closure and reclamation plan has three objectives that provide assurance to the Province that the site will be left in a condition that will limit the future liability to the people of BC:

 

to provide stable landforms

 

to re-establish productive land use

 

to protect terrestrial and aquatic resources.

 

Provision of Stable Landforms

 

The design of KSM’s permanent mine-related landforms, such as the open pits, the TMF, the WSD that will impound the WSF at the mine site, seepage containment dams, and the RSF, has been undertaken to ensure long-term stability during mine operations, after mine closure, and after reclamation works are complete.

 

Re-establishment of Productive Land Use

 

The pre-development land use and conditions form the basis for setting the end land use and capability objectives. The goal is to return the site to a use consistent with the current land uses. The current land use information has been obtained from the environmental and socio-economic baseline studies (Appendix 23-A of the Application/EIS) (Rescan 2013). These studies were undertaken in consultation with the KSM Project EA Working Group which includes provincial and federal government agencies, Nisga’a Nation, Tahltan Nation, Gitxsan Nation, Skii km Lax Ha, and the Gitanyow First Nation.

 

The end land use objective will be primarily to provide for wildlife habitat for the described wildlife species, including bears, mountain goats, and moose.

 

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20.6.2Soil Handling Plan

 

The general goal of reclamation is to restore, where possible, the equivalent land capability so that end land use objectives can be achieved. Site reclamation planning will include the conservation of soil materials suitable for reclamation purposes in areas disturbed by mining, and these areas will be re-vegetated, where feasible. The landforms resulting from KSM mine development will also be designed, where possible, and reclaimed to accommodate the desired end land use objective, involving development of appropriate and functional ecosystems, as supported by appropriate soil material handling and re-vegetation strategies.

 

20.6.3Closure and Reclamation Planning

 

Mitchell/East Mitchell Pit Closure

 

In the 2022 PFS, the Mitchell deposit is proposed to be mined as an open pit from Year -2 to Year 24, with an ultimate wall height of 1,230 m. Closure of the Mitchell pit includes backfilling with water to form a pit lake and placing large rocks on the benches to discourage wildlife access to the pit lake. The Mitchell pit cannot be backfilled with water until underground mining is completed. The East Mitchell pit receives waste from the final mining phase of the Mitchell pit and the East Mitchell pit will be graded to allow positive drainage to the Mitchell pit lake.

 

Sulphurets Pit Closure

 

The Sulphurets pit will be backfilled with the waste rock from the Mitchell pit to a level that creates positive drainage to the WSF.

 

Mitchell OPC

 

Closure

 

At the completion of mining, the Mitchell OPC will be decommissioned. Equipment will be removed from the site. The electrical substation will remain. All other structures will be dismantled and removed. Foundations will be broken up, and the concrete rubble will be buried on site. Any soils that are contaminated with fuel will be excavated and treated at a landfarm to remediate the soil.

 

Reclamation

 

The Mitchell OPC ground surface will be ripped and covered with crushed rock and up to 50 cm of topsoil prior to re-vegetation with native species.

 

McTagg Diversion Tunnels Closure

 

The McTagg Power Plant will be located east of the Gingras Creek bridge and will be maintained indefinitely to generate electricity. The electricity will be used to operate the HDS WTP, or sold for use in the provincial electricity grid.

 

Water Storage Facility and Water Treatment Plant

 

The WSF will remain in service after mine closure to continue collecting contact water that requires treatment.

 

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The HDS WTP and support infrastructure will remain in operation during the closure and post-closure phases. The plant will operate primarily in the spring, summer, and fall months, and minimally in the winter. The lime material will be transported to the site and consumed during these warmer periods. At closure and post-closure, the filter cake (sludge) will be hauled by truck during the summer to the top of the RSF and placed in an engineered landfill.

 

An ion exchange Selenium WTP located near the WSD will remain in service after mine closure, until receiving environment water quality standards can be met without treatment.

 

At closure and post-closure, the tunnels will be required to provide ongoing access to the mine site because the CCAR, which serves the mine site during operation, will be decommissioned. As the HDS WTP will continue to operate post-closure, lime will be required and will be transported from the PTMA through the tunnels to the mine site. All supplies for monitoring, maintenance, and the operation of the HDS WTP will be transported through the tunnels. The train logistics system will be reconfigured and simplified to handle this much reduced traffic requirement.

 

Treaty Process Plant, Carbon-in-Leach Plant, and Other Structures Closure

 

Closure

 

Several structures at the PTMA that will be closed at the completion of mining including the Treaty Process Plant, the CIL Plant, the Treaty OPC waste management facilities, the Treaty OPC Batch Plant, the crusher building, and several other structures such as the warehouse and lab. All of these structures contain equipment that must be removed at closure.

 

Once all of the equipment has been removed, the buildings will be dismantled and the materials will be moved off-site for recycling or disposal. Any contaminated soil will be collected and placed in the landfarm. Any materials that can be incinerated will be incinerated on site. The concrete foundations of the various buildings will be broken up, buried, and used for road maintenance or as armouring in TMF reclamation. All equipment and debris will be removed from around the structures.

 

Reclamation

 

The footprint areas and the area surrounding the buildings will be deep-ripped to reduce compaction and to improve surface drainage. Approximately 30 cm of soil will be spread over the area as the surface material to improve conditions for vegetation establishment and growth of native plants.

 

Coarse and Fine Ore Stockpiles Closure

 

Closure

 

All of the ore in the coarse and fine stockpiles will be processed. The footprint areas will be cleaned and ripped to reduce compaction and to increase downward drainage.

 

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Reclamation

 

The footprint areas will be covered with a lime mixture and topsoil, and then vegetated with native plants as described for the RSF.

 

Treaty Ore Preparation Complex Batch Plant Stockpile Closure

 

The Treaty OPC may contain stockpiled materials left over from operations consisting of sands and gravels. Any remaining materials will be used for road maintenance or as rip-rap along the beach edges in the TMF.

 

Structures Required Post-closure

 

Some structures in the PTMA will be required for on going monitoring of mine post-closure. These include:

 

the office complex

 

the ambulance building

 

substation 1

 

the operating camp incinerator

 

the administration building

 

the operating camp (reduced in size from the operation phase).

 

As described above, the MTT will remain open because the HDS WTP will continue to operate post-closure, and reagents, including lime and personnel operating the HDS WTP, will be transported through the tunnels. A smaller camp will be required to accommodate personnel.

 

Tailing Management Facility Closure

 

Following operations, the TMF will be reclaimed to provide for wildlife and wetland habitat. The dams and beaches of the TMF will be reclaimed in stages, with the North Cell being reclaimed during operation, the South Cell during closure, and the Centre Cell during post-closure. The TMF facility will be reclaimed in accordance with the described reclamation and closure plan as follows.

 

The TMF North Cell will be closed approximately 5 years after tailing deposition into it ceases following expansion of the beach area and the time for water quality to improve for discharge.

 

The South Cell will be closed in a similar manner to the North Cell.

 

The Centre Cell is the last cell to be closed, which will be about 5 years following mine closure. The CIL tailing in the CIL lined pond will be covered with approximately 1 m of non-reactive tailing and submerged under 5 m of water.

 

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Access Roads

 

Treaty Creek Access Road Closure

 

During closure and post-closure, the TCAR will provide the only remaining road access to the Property. All materials and personnel will be transported via the TCAR, which will extend to the Treaty OPC and the MTT portals, which will all remain open to allow for access to the mine site post-closure.

 

Coulter Creek Access Road Closure

 

The CCAR will be decommissioned post-closure. The bridges will be dismantled, and materials that are combustible will be burned. Concrete will be broken and used as rip-rap along the creeks, if required, to reduce potential surface erosion that could occur during the dismantling of the bridges.

 

Culverts will be removed to restore natural drainage patterns. Cross-ditching will provide drainage across roads and will reduce the potential for surface erosion. The surface of the road and any compacted areas will be ripped, where required, to promote surface drainage and to reduce runoff and potential road bed failure.

 

Quarry and Borrow Sources Closure

 

The borrow areas will be cleared and grubbed. The quarry and borrow areas will be re-sloped and re-contoured to ensure escape routes for wildlife and to restore natural landscape.

 

Mini Hydro Plant and Energy Recovery

 

Energy recovery and a mini-hydro plant at the end of the McTagg Diversion Tunnel (MTDT) is included in the KSM mine development plan. This plant will generate electrical power by making use of facilities already included in the 2022 PFS, resulting in significant net energy savings. The power plant will be used to provide electricity for the mine closure requirements, or the electricity may be sold back to BC Hydro under the Standing Offer Program.

 

Water Treatment Plant Energy Recovery

 

Water pumped from the water storage pond to the HDS WTP will generate electric power. A small impulse Turgo-type turbine will be used. The output may be fed into the plant power distribution system at the HDS WTP. This facility will continue to operate after mine closure.

 

Operation Camps

 

Closure

 

The Mitchell Operating Camp and the Treaty Operating Camp will generally include portable trailers, an incinerator, materials and equipment storage areas, a helicopter pad, a helicopter fuelling area, fuel storage, a septic field, water/sewage treatment, and diesel generators. The portables will be set up so that they can be dismantled and used at the different sites, as required.

 

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The operation camp sites will be reclaimed to a slope compatible with the surrounding natural topography. Any contaminated soils will be relocated to the landfarm. The high traffic areas will be ripped in two directions to increase surface drainage and to allow for deeper root penetration. All construction camps will be decommissioned and reclaimed during early operations.

 

Reclamation

 

Prior to construction, topsoil will have been salvaged from the camp site areas and stockpiled along the edges of the camps. At closure, this soil will be spread over the disturbed areas. These areas will then be re-vegetated with the native grasses, shrubs, and tree seedlings that were described for the RSF.

 

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21.0Capital and Operating Cost Estimates

 

21.1Initial Capital Costs

 

An initial capital cost of US$6.432 billion was estimated for the 2022 PFS, based on capital cost estimates developed by the following consultants:

 

MMTS: open pit mining, mine roads, ore trains, and infrastructure and water diversion tunnels

 

Kamburt Civil Consulting Ltd. (KCC): costing of civil earthworks related to the TMF, designed by KCB

 

EBC Inc. (EBC): costing of the WSD, designed by KCB

 

Tetra Tech: process plant and associated infrastructure, including plant site preparation, water treatment plant, construction camps and main access roads

 

Brazier: permanent power supply, fire detection, mini hydro plant, and energy recovery systems

 

BGC: landslide management, avalanche management, and pit depressurization

 

Seabridge: Owner’s and fish compensation costs.

 

All currencies in this section are expressed in US dollars, unless otherwise stated. Costs have been converted using a fixed currency exchange rate of US$0.77 to Cdn$1.00.

 

The expected accuracy range of the capital cost estimate is +25/-10%.

 

The costs stated in Table 21.1 include only initial capital, which is defined as all costs to build the facilities that mine, transport, and process ore to produce first concentrate and doré. Costs incurred during ramp-up of the mine and process plant in Year 1, through commercial production, are included in the operating costs in Section 21.3.

 

This estimate was prepared with a base date of Q1/Q2 2022. The estimate does not include any escalation past this date. Budget quotations were obtained for major equipment; vendors provided equipment prices, delivery lead times, spare allowances, and freight costs to a designated marshalling yard in northern BC, with some exceptions for delivery points to different BC locales. The quotations used in this estimate were obtained in Q1/Q2 2022, and are budgetary and non-binding.

 

For non-major equipment, costing is based on in-house data, quotes from previous studies. No cost escalation is included.

 

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All equipment and material costs include Incoterms FCA. Other costs such as spares, taxes, duties, freight, and packaging are covered separately in the estimate as indirect costs.

 

The initial capital cost summary and its cost breakdown structure (CBS), which is based on the work breakdown structure (WBS) for the 2022 PFS, are presented in Table 21.1.

 

Table 21.1 Initial Capital Cost Summary

 

Major
Area
No.
Major Area Description Cost
(US$ M)
1 – Direct Costs
1.1 Mine Site 1,420
1.2 Process 2,004
1.3 TMF 513
1.4 Environmental 15
1.5 On-site Infrastructure 39
1.6 Off-site Infrastructure 76
1.7 Permanent Electrical Power Supply and Energy Recovery 121
Total Direct Costs 4,188
2 – Indirect Costs
2.91 Construction Indirect Costs 565
2.92 Spares 55
2.93 Initial Fills 26
2.94 Freight and Logistics 110
2.95 Commissioning and Start-up 7
2.96 EPCM 299
2.97 Vendor’s Assistance 29
Total Indirect Costs 1,091
3 – Owner’s Costs
3.98 Owner’s Costs 204
4 – Contingency
4.99 Contingency 949
2022 PFS Capital Cost Total 6,432
Note:Costs have been rounded to the nearest million dollars.

 

21.1.1Exclusions

 

The following items are not included in the capital cost estimate:

 

substantial start (early works to develop site access and construct permanent infrastructure)

 

force majeure

 

schedule delays, such as those caused by:

 

qmajor scope changes

 

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qunidentified ground conditions

 

qlabour disputes

 

qenvironmental permitting activities

 

qabnormally adverse weather conditions

 

receipt of information beyond the control of the EPCM contractors

 

salvage value for assets only used during construction

 

cost of financing (including interests incurred during construction)

 

sales taxes (PST, GST and HST)

 

royalties or permitting costs, except as expressly defined

 

schedule acceleration costs

 

forward inflation

 

abnormal price fluctuations due to pandemic, geopolitical instability or interruptions in logistics, supply chain and world trade. working capital

 

cost of this study and future feasibility study

 

escalations beyond effective date of this study

 

growth factors in design and engineering

 

uncertainties in geotechnical or hydrogeological conditions

 

sunk costs.

 

Labour Rates

 

A standard labour rate has been applied to various areas of the 2022 PFS. The standard construction labour rate used is US$83.93/h (Cdn$109.00/h) and is considered fully burdened. The base labour rate of US$34.50 (Cdn$44.81) was calculated from a combination of union rates and current labour rates at operating mines in BC published by the unions and BC Labour Relations Board, and across all construction crafts.

 

21.1.2Direct Costs

 

Mine Site

 

The Mine Site initial capital costs totals US$1.420 billion as presented in Table 21.2.

 

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Table 21.2 Mine Site Capital Costs

 

Area
No.
Area Description Cost
(US$ M)
1.1.01 Open Pit Mining 635
1.1.02 WSF 289
1.1.03 Surface Water Management Including Water Diversion Tunnels) 275
1.1.04 Water Treatment 110
1.1.05 Ancillary Buildings 11
1.1.06 Site Services and Utilities 6
1.1.07 Power Supply and Distribution 8
1.1.08 Camps 53
1.1.13 Geohazards 34
Mine Site Capital Costs Total 1,420
Note:Costs have been rounded to the nearest million dollars.

 

Open Pit Mining

 

Open pit capital costs were derived from a combination of supplier quotes and historical data collected by MMTS. Pre-production operating costs of the mining fleet and earth works for pioneering and construction quarries are included in the open pit mining capital.

 

Water Storage Facility and Surface Water Management

 

The WSF capital cost was estimated at US$288.5 million. The main dam structure is the largest cost at US$148.5 million. The capital cost of surface water management, including water diversion tunnels (MDT and MTDT), is US$275 million.

 

Process

 

The battery limits for process are the Mitchell OPC (primary crusher), the MTT (including Saddle), trains, and the Treaty OPC (crushing, grinding, flotation and leaching equipment; the concentrator and other Treaty OPC buildings). TMF costs are separate from process costs. Process capital costs were estimated at US$2,004 million.

 

TMF

 

TMF capital costs were estimated at US$513.3 million. Initial capital comprises initial TMF perimeter diversions; the North Dam, Splitter Dam, and Saddle Dam; and associated seepage collection dams that form the North and Centre cells. This includes basin preparation for the starter basins for the North Cell and Centre Cell, preparing and lining the starter basins, and provision of liner drainage. Details of process and OPC capital costs are presented in Table 21.3.

 

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Table 21.3 Process-Treaty OPC Capital Cost Estimate

 

Area
No.
Area Description Cost
(US$ M)
1.2.01 Primary Crushing 56
1.2.02 Ore Delivery Tunnel 649
1.2.05 MTT Material Transportation 307
1.2.06/07 TWTP– Saddle & Treaty 8
1.2.08 Treaty OPC – Coarse Ore Stockpile 137
1.2.09 Treaty OPC – Secondary Crushing 100
1.2.11 Treaty OPC – Tertiary Crushing - HPGR 102
1.2.12 Process Building 106
1.2.13 Primary Grinding 102
1.2.14/15 Flotation 75
1.2.16 Pyrite Concentrate Regrinding 29
1.2.17 Cyanide Leaching 74
1.2.18 Gold/Silver Refinery 14
1.2.19 Copper Concentrate Handling 13
1.2.20/21 Molybdenum Flotation, Leaching & Concentrate Handling’ 10
1.2.22 Cyanide Recovery and Destruction 35
1.2.23 Reagent Area 7
1.2.24 Plant Control System 5
1.2.25 Site Services and Utilities 26
1.2.27/30 Treaty Temporary Laydown & Ancillary Buildings 39
1.2.31 Process Plant Utilities and Infrastructure 21
1.2.32 Treaty OPC – Mobile Equipment 11
1.2.33 Power Supply and Distribution 7
1.2.34 Treaty OPC – Roads 3
1.2.35 Treaty Operations and Construction Camps 67
Process-Treaty OPC Capital Cost Total 2,004
Note:Costs have been rounded to the nearest million dollars.

 

21.1.3Indirect Costs

 

Indirect costs for the initial capital cost estimate were estimated at US$1,091 million, as shown in Table 21.4.

 

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Table 21.4 Indirect Capital Costs

 

Area
No.
Area Description Cost
(US$ M)
2.91 Construction Indirect Costs 565
2.92 Spares 55
2.93 Initial Fills 26
2.94 Freight and Logistics 110
2.95 Commissioning and Start-up 7
2.96 EPCM 299
2.97 Vendor’s Assistance 29
Indirect Capital Costs Total 1,091
Note:Costs have been rounded to the nearest million dollars.

 

21.1.4Owner’s Costs

 

An estimate of US$204 million is included in the initial capital cost estimate for Owner’s costs. This cost had been calculated by Seabridge from first principals based on personnel requirements and onboarding of Owner’s staff for supporting both the latter part of the commissioning effort for construction and onboarding to fill all operations and G&A departments. In addition to labour costs. Owner’s costs include IROCs, off-site office staffing, off-site office facilities, travel, off-site office general expenses, recruiting and training expenses, consulting, insurance, general field expenses, and mineral lease/claims costs. Environmental department in the Owner’s costs include environmental monitoring programs, community relations, communication and public relations, wetland compensation, and permitting costs.

 

21.1.5Contingency

 

A contingency allowance is included to cover additional costs that could occur as a result of more detailed design or unexpected site conditions. The estimated contingency cost is US$949 million. The contingency estimate was developed on a line-item basis to account for the specific design details and information available for each area, rather than a single value applied to the sum of all directs and indirects. The values applied range from 9% to 25%.

 

21.2Sustaining Capital Costs

 

The sustaining capital costs are all capital costs required from Year 1 of operations to sustain the mining operation for the LOM. Details of the total sustaining cost of US$3.210 billion, required for the LOM, is presented in Table 21.5.

 

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Table 21.5 Sustaining Capital Costs

 

Major Area
No.
Major Area Description Cost
(US$ M)
1.1 Mine Site 1,766
1.2 Process (includes process buildings and utilities) 309
1.3 TMF (incl. East Catchment Tunnel) 630
1.4 Environmental 8
1.6 Off-site Infrastructure 11
1.7 Permanent Electrical Power Supply and Energy Recovery 46
2.91 Construction Indirects 49
2.92 Spares 23
2.94 Freight and Logistics 7
2.95 Vendor’s Assistance, Commissioning & Start-up 18
2.96 EPCM 1
4.99 Contingency 343
Sustaining Capital Cost Total 3,210
Note:1. Costs have been rounded to the nearest million dollars.
  2. Mine Site costs include surface water management and diversion tunnels.
  3. Process costs include MTT and ore transport train system costs

 

21.2.1Mine Site

 

The sustaining capital for the Mine Site is US$1.766 billion presented in Table 21.6. This number covers the direct capital costs for the LOM and includes all open pit and underground mining operations, as well as the Selenium WTP and the geohazards direct capital cost.

 

Table 21.6 Mine Site Sustaining Capital Costs

 

Area
No.
Area Description Cost
(US$ M)
1.1.01 Open Pit 1,454
1.1.03 Surface Water Management 114
1.1.04 Water Treatment 120
1.1.05 Ancillary Buildings 47
1.1.12 Sulphurets Pit Power Supply 2
1.1.13 Geohazards 29
Mine Site Sustaining Capital Cost Total 1,766
Note:Costs have been rounded to the nearest million dollars.

 

21.2.2Open Pit Mining

 

Sustaining capital is based on both fleet expansions and unit replacements over the LOM. Major fleet expansions were planned for Years 1 and 24 as the mining rate and haul distances increase. Capital replacement costs for mobile equipment were calculated based on the expected life of the equipment, the cost of the unit, and the utilization for that equipment.

 

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21.2.3Mine Site Water Treatment

 

HDS WTP

 

The sustaining water treatment cost for the Mine Site is US$120 million.

 

The water treatment capacity at the HDS WTP will increase with the addition of one more clarifiers to cater for the increased water flow that is expected.

 

When the HDS WTP is commissioned in Year -2, the initial maximum throughput capacity will be 3.0 m3/s.

 

In Year 1, the plant capacity will increase to 4.0 m3/s final capacity. One additional circuit will be constructed and operated in parallel to the existing three circuits.

 

Selenium WTP

 

The sustaining capital cost of the Se WTP is US$134 million.

 

In Year 5, a 500 L/s Selenium WTP, located adjacent to the WSF near the toe of the Mitchell RSF, will be constructed and become operational to treat seepage from select pit sources, seepage from the RSFs, and water pumped from the WSF.

 

21.2.4Process

 

Process sustaining capital costs include the Sulphurets pit primary crusher in Year 1, and the expansion of the process plant that is designed to increase throughput to 195,000 tpd by Year 3.

 

Additional trains are included for the increase in MTT material transport and for the decline in mechanical availability over time.

 

The estimate also includes a replacement allowance for major process equipment.

 

The process sustaining capital cost was estimated at US$309 million.

 

21.2.5Tailing Management Facility

 

The sustaining capital for the TMF is US$630 million.

 

Dam raising for the North and Centre cells is accounted for in both sustaining capital and operating expenses. Borrow, haul, and placement of till core and expansion of basin preparation, drains and CIL Residue Cell liner are considered sustaining capital that begin in Year 1. The processing and placement of cycloned sand is considered an annual operating expense for all stages through the LOM.

 

The South Cell includes additional sustaining capital items occurring between Year 16 and the LOM. These include development of additional perimeter diversions for the South Cell and the Southeast Starter Dam and associated seepage dam.

 

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Ancillary expenses such as seepage pumping and monitoring are included as operating expenses throughout the LOM.

 

21.2.6Other Sustaining Capital Costs

 

Other sustaining capital costs were estimated at US$57 million, mostly consisting of fish habitat compensation, site access road and permanent power supply. The indirect sustaining capital costs were estimated at US$97 million and sustaining capital contingency costs were estimated at US$343 million.

 

21.3Operating Costs

 

The average operating cost was estimated at US$11.20/t milled at the final nameplate process rate of 195,000 t/d, or US$11.36/t for the LOM average. The operating cost estimate accounts for the energy recovery credits (approximately US$0.07/t milled LOM) from mini-hydro power stations and the cost estimated for PST (approximately US$0.13/t milled LOM).

 

The mining operating costs are LOM average unit costs calculated by dividing the total LOM operating costs by LOM milled tonnages. The costs exclude mine pre-production costs.

 

The cost distribution for each area is shown in Figure 21.1, excluding the energy recovery credits from the recovered hydro-energy during mining operations and the applicable PST.

 

The operating cost estimates in this section are based on budget prices obtained in Q1/Q2 2022 and/or from databases of the consulting firms involved in preparing the operating cost estimates.

 

When required, certain costs in this report have been converted using a fixed currency exchange rate of Cdn$1.00 to US$0.77. The expected accuracy range of the operating cost estimate is +25/-10%. A summary of the operating costs is presented in Table 21.7.

 

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Table 21.7 Operating Cost Summary

 

  At the Nominal Feed
Rate of 195,000 t/d*
LOM
Average
(US$/t
milled)
(US$ M/a) (US$/t milled)
Mine
Mining Costs – Mill Feed 235.6 3.31 3.31
Mill
Process 437.6 6.15 6.31
G&A and Site Service
G&A 51.2 0.72 0.75
Site Service 22.0 0.31 0.31
TMF and SWM
TMF Dam Management 7.5 0.11 0.11
Selenium Water Treatment 19.6 0.28 0.25
HDS Water Treatment 15.3 0.22 0.22
Mine Site Water Pumping 2.2 0.03 0.03
Energy Recovery -4.8 -0.07 -0.07
Provincial Sales Tax 9.6 0.14 0.14
Total Operating Cost 795.8 11.20 11.36
Note:The nominal feed rate estimate excludes mine operating costs and is based on a mill feed rate of 195,000 t/d; the costs do not reflect higher unit costs late in the mine life when the mill feed rates are lower. Costs have been rounded to the nearest hundred thousands of dollars.

 

Figure 21.1 Operating Cost Distribution

 

Source: Tetra Tech, 2022

 

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Sustaining capital costs including all capital expenditures after process plant first production are excluded from the operating cost estimate.

 

Mining Personnel

 

Salaries for the supervisory and administrative job categories, and all hourly employee labour rates, are based on an industry-focused and location-specific labour survey conducted by Seabridge. Burdens are included in the salaries and labour rates. The payments include base salaries/labour rates, holiday and vacation pay, government prescriptive benefits (e.g., Canadian Pension Plan, workman compensation insurance, etc.), discretionary employer sponsored benefits, and tool allowance costs.

 

Labour factors in man-hours/equipment operating hours were estimated for operations and maintenance labour for each of the equipment types. Labour costs were calculated by multiplying the labour factor by the equipment operating hours, and labour costs were allocated to the equipment where labour had been assigned. The total hours required for each job type on all the equipment were added, and any additional labour required to complete a crew was assigned to an unallocated labour category.

 

Process Personnel

 

Salary/wage rates for management, technical support and operation are based on an industry-focused and location-specific labour survey conducted by Seabridge. Burdens are included in the salaries and labour rates. The payments include base salaries/labour rates, holiday and vacation pay, government prescriptive benefits (e.g., Canadian Pension Plan, workman compensation insurance, etc.), discretionary employer sponsored benefits, and tool allowance costs.

 

21.3.1Open Pit Mine Operating Costs

 

Open pit mine operating costs, including operating and maintenance salaried staff and hourly labour, equipment major component and running repairs, fuel, power (excluding trolley power for the trucks),, and all other consumable goods, were derived from a combination of supplier quotes and historical data collected by MMTS. The quantities of consumables required were determined for each specific open pit mining activity from vendor input and in-house experience. Labour factors for operations and maintenance of the open pit mining equipment was also estimated based on vendor input and MMTS experience. These inputs were used to build up open pit mine operating costs from first principles.

 

Freight costs for all consumable goods and fuel are included in the estimate as part of the budgetary quotations.

 

Major component replacement for larger pieces of mobile equipment were calculated based on the expected life of the major component, the cost of the component, and the fleet size for that equipment. This puts large component repair costs into future years, giving a more representative LOM cash flow.

 

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The cost of minor parts and running repairs were estimated as an hourly operating cost for the mining equipment. The cost of geotechnical support works for mine operations is included in the mining operating costs.

 

GME is a category for open pit mine operations, mine maintenance, and technical services departmental overhead costs. It consists of costs for all salaried supervisory and technical staff, a consumable and rental allowance, crane rentals, and software and fleet management systems’ licensing and maintenance. This category is a fixed cost, and does not vary by production or fleet size, with the exception of ramp-ups to full staffing.

 

The distribution of unit cost by mining area is shown in Figure 21.2.

 

Figure 21.2 LOM Average Unit Operating Cost for Open Pit Mining (US$/t Material Mined) – excludes mine pre-production costs

 

Source: Tetra Tech, 2022

 

21.3.2Process Operating Costs

 

Summary

 

The LOM average annual process operating costs for the different mineralizations were estimated as:

 

Mitchell mineralization: US$433 million (US$6.08/t milled)

 

East Mitchell mineralization: US$440 million (US$6.19/t milled)

 

Sulphurets mineralization: US$455 million (US$6.39/t milled).

 

The process operating costs for these mineralizations are based on a process rate of 195,000 t/d and 94% plant availability. The estimated average operating cost for the Sulphurets ore is higher than the ores from the Mitchell deposits, mainly due to harder mineralization for the Sulphurets ore, as compared to the Mitchell deposits. Due to the variations in operating costs for the different deposit ores, the average operating costs were estimated based on the ratio of the different ore tonnages processed and their individual operating costs.

 

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The estimated process operating costs are summarized in Table 21.8 and Table 21.9, which include:

 

personnel requirements, including supervision, operation and maintenance; salary/wage levels based on an industry-focused and location-specific labour survey conducted by Seabridge, the payments include base salaries/labour rates and various burdens

 

liner and grinding media consumption estimated from the Bond ball mill work index and abrasion index equations; maintenance supplies are based on approximately 6% of major equipment capital costs

 

reagents based on metallurgical test results

 

Ore transport train maintenance

 

other operation consumables including laboratory, filtering cloth, and service vehicles consumables

 

power consumption for the process plant

 

no taxes or import duties are included in the process operating cost estimate, unless specified.

 

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Table 21.8 Summary of Process Operating Costs by Deposit

 

Area  Personnel Mitchell Sulphurets East Mitchell

Annual Cost

(US$ M)

Unit Cost

(US$/t

milled)

Annual Cost

(US$ M)

Unit Cost

(US$/t

milled)

Annual Cost

(US$ M)

Unit Cost

(US$/t

milled)

Human Power
Operating Staff 43 5.3 0.07 5.3 0.07 5.3 0.07
Operating Labour 152 13.3 0.19 13.3 0.19 13.3 0.19
Maintenance 87 8.2 0.12 8.2 0.12 8.2 0.12
Subtotal Human Power 282 26.7 0.38 26.7 0.38 26.7 0.38
Major Consumables and Supplies
Major Consumables
Metal Consumables   118.4 1.66 135.7 1.91 125.2 1.76
Reagent Consumables   178.8 2.51 178.8 2.51 178.8 2.51
Supplies
Maintenance Supplies   27.5 0.39 27.5 0.39 27.5 0.39
Operating Supplies   3.2 0.04 3.2 0.04 3.2 0.04
Subtotal Consumable and Supplies 327.8 4.61 345.1 4.85 334.6 4.70
Power Supply   78.5 1.10 83.2 1.17 79.1 1.11
Subtotal Power   78.5 1.10 83.2 1.17 79.1 1.11
Process Operating Cost Total 282 433.0 6.08 455.1 6.39 440.4 6.19
Note:Sums may not add due to rounding.

 

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Table 21.9 Operating Costs per Area of Operation by Deposit

 

Area Personnel Mitchell Sulphurets East Mitchell

Annual Cost

(US$ M)

Unit Cost

(US$/t

milled)

Annual Cost

(US$ M)

Unit Cost

(US$/t

milled)

Annual Cost

(US$ M)

Unit Cost

(US$/t

milled)

Crushing, Grinding and Copper Flotation Plant 156 290.7 4.08 312.6 4.39 298.1 4.19
Tunnel Transport 43 13.8 0.19 13.8 0.19 13.8 0.19
Molybdenum Flotation Plant 8 7.6 0.11 7.6 0.11 7.5 0.11
Leach Plant 43 49.9 0.70 49.9 0.70 49.9 0.70
Cyanide Solution/Residue Handling 12 62.3 0.88 62.3 0.88 62.3 0.88
Tailing Pumping/Reclaim Water 20 8.8 0.12 8.8 0.12 8.8 0.12
Total 282 433.0 6.08 455.1 6.39 440.4 6.19
Note:Sums may not add due to rounding.

 

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21.3.3TMF Dam Management Operating Costs

 

On average, the operating costs for ongoing tailings dam construction by cycloning sand, starting from the secondary cyclone station and including seepage water pumping costs, were estimated to be approximately US$7.7 million/a, or US$0.11/t milled. The other tailings facility construction related costs are included in sustaining capital costs.

 

21.3.4Mine Site Water Management Costs

 

Overall SWM at the Mine Site includes HDS site water treatment, selenium removal treatment, and site water pumping. The average LOM annual cost for Mine Site SWM was estimated to be approximately US$42.6 million/a, or US$0.52/t milled, at a mill feed rate of 195,000 t/d.

 

The estimated average LOM operating cost for the HDS WTP is approximately US$0.22/t milled, at a mill feed rate of 195,000 t/d, or US$0.31/m3 water, treated at an average flow rate of approximately 49.8 Mm3/a. The maintenance manpower will come from the overall mine site maintenance team. The major cost for HDS water treatment is reagent consumption at US$0.14/t milled.

 

The estimated LOM average operating cost for the Selenium WTP is approximately US$0.27/t milled at a mill feed rate of 195,000 t/d, or US$1.24/m3 water, treated at an average flowrate of approximately 15.8 Mm3/a. The maintenance manpower will come from the overall site services maintenance team. The major cost for selenium water treatment is US$0.11/t milled for reagent consumption and maintenance. Power consumption was estimated to be approximately 25.4 GWh/a.

 

21.3.5General and Administrative

 

G&A costs are costs that do not relate directly to mining or processing operating costs. These costs include:

 

personnel: executive management, staffing in accounting, supply chain and logistics, human resources, external affairs functions, and other G&A departments

 

expenses : including insurance, off site offices, administrative supplies, medical services, legal services, human resource related expenses, travelling, community and environmental programs, accommodation/camp costs, air/bus crew transportation, regional and property taxes, and external assay/testing.

 

The G&A costs were estimated at approximately US$39.5 million/a, or US$0.56/t milled at a nominal mill feed rate of 195,000 t/d, including approximately US$0.25/t for personnel and US$0.47/t for general expenses. The major costs are accommodation and crew air transportation, estimated at about US$14.5 million/a.

 

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21.3.6Site Services

 

The overall site service cost was estimated at US$0.31/t milled or approximately US$22.0 million/a. The estimate is based on requirements for this remote site in northern BC and on in-house experience. The estimate includes the following:

 

personnel: general site service human power

 

site mobile equipment and light vehicle operations

 

potable water and waste management

 

general maintenance for yards, roads, fences, and buildings

 

off site operation expense

 

building heating

 

power supply

 

avalanche control.

 

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22.0Economic Analysis

 

22.1Introduction

 

Tetra Tech prepared an economic evaluation of the 2022 PFS based on both a pre-tax and a post-tax basis in a single financial model. For the 33-year LOM and 2.292 billion tonne Mineral Reserve, the following pre-tax financial results were calculated using the 2022 Base Case metal prices:

 

20.1% IRR

 

3.4-year payback on US$6.4 billion initial capital

 

US$13.5 billion NPV at a 5% discount rate.

 

Seabridge engaged PwC in Toronto, Ontario to review the tax component of the model for the post-tax economic evaluation for this 2022 PFS with the inclusion of applicable income and mining taxes. PwC is an Ontario limited liability partnership, which is a member firm of PricewaterhouseCoopers International Limited, each member firm of which is a separate legal entity.

 

The following post-tax financial results were calculated:

 

16.1% IRR

 

3.7-year payback on US$6.4 billion initial capital

 

US$7.9 billion NPV at a 5% discount rate.

 

The 2022 Base Case results apply the following key inputs:

 

gold – US$1,742/oz

 

copper – US$3.53/lb

 

silver – US$21.90/oz

 

molybdenum – US$18.00/lb

 

exchange rate – Cdn$1.00 to US$0.77.

 

Sensitivity analyses, along with multiple additional metal price scenarios, were developed to evaluate the 2022 PFS economics.

 

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22.2Forward-looking Statements

 

This document contains “forward-looking information” within the meaning of Canadian securities legislation and “forward-looking statements” within the meaning of the United States Private Securities Litigation Reform Act of 1995. This information and these statements, referred to herein as “forward-looking statements” are made as of the date of this document. Forward-looking statements relate to future events or future performance and reflect current estimates, predictions, expectations or beliefs regarding future events and include, but are not limited to, statements with respect to: (i) the estimated amount and grade of Mineral Reserves and Mineral Resources; (ii) estimates of the capital costs of constructing mine facilities and bringing a mine into production, of operating the mine, of sustaining capital and the duration of payback periods; (iii) the estimated amount of future production, both material processed and metal recovered; (iv) estimates of operating costs, life of mine costs, net cash flow, net present value (NPV) and economic returns from an operating mine; and (vi) the assumptions on which the various estimates are made being reasonable.

 

All forward-looking statements are based on the author’s’ current beliefs as well as various assumptions made by them and information currently available to them. These assumptions are set forth throughout this Report, and some of the principal assumptions include: (i) the presence of and continuity of metals at estimated grades; (ii) the geotechnical and metallurgical characteristics of rock conforming to sampled results; (iii) the quantities of water and the quality of the water that must be diverted or treated during mining operations; (iv) the capacities and durability of various machinery and equipment; (v) anticipated mining losses and dilution; (vi) metallurgical performance; and (vii) reasonable contingency amounts. Although the QPs consider these assumptions to be reasonable based on information currently available to them, they may prove to be incorrect. Many forward-looking statements are made assuming the correctness of other forward-looking statements, such as statements of net present value and internal rates of return, which are based on most of the other forward-looking statements and assumptions herein.

 

By their very nature, forward-looking statements involve inherent risks and uncertainties, both general and specific, and risks exist that estimates, forecasts, projections and other forward-looking statements will not be achieved or that assumptions do not reflect future experience.

 

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22.3Pre-Tax Model

 

Metal revenues projected in the 2022 PFS cash flow models are based on the average metal values indicated in Table 22.1.

 

Table 22.1 Metal Production from the KSM Mine

 

  Years 1 to 7 LOM
Total Tonnes to Mill (Mt) 441 2,292
Annual Average Tonnes to Mill (Mt) 63 69
Average Grades
Gold (g/t) 0.89 0.64
Copper (%) 0.21 0.14
Silver (g/t) 3.0 2.2
Molybdenum (ppm) 52 76
Total Production
Gold (Moz) 9.89 33.90
Copper (Mlb) 1,756 5,872
Silver (Moz) 26.97 97.89
Molybdenum (Mlb) 14.64 140.20
Average Annual Production
Gold (Moz) 1.413 1.027
Copper (Mlb) 250.8 178.0
Silver (Moz) 3,853 2,966
Molybdenum (Mlb) 2.09 4.25

 

22.3.1Financial Evaluations: NPV and IRR

 

The production schedule has been incorporated into the 100% equity financial model to develop annual recovered metal production from the relationships of tonnage processed, head grades, and recoveries.

 

Metal revenues, principally gold and copper, were calculated based on each scenario’s prices. Operating cost for mining, processing, site services, G&A, tailing storage and handling and water treatment, energy recovery areas, as well as off-site charges (smelting, refining, transportation, and royalties) were deducted from the revenues to derive annual operating cash flow.

 

Initial and sustaining capital costs as well as closure and reclamation costs have been incorporated on an annual basis over the mine life and deducted from the operating cash flow to determine the net cash flow before taxes. Initial capital expenditures include costs accumulated prior to first production of concentrate, including all pre-production mining costs. Sustaining capital includes expenditures for mining and processing additions, replacement of equipment, and TMF expansions.

 

Initial and sustaining capital costs applied in the economic analysis are US$6.4 billion and US$3.2 billion, respectively. LOM PST applicable to initial and sustaining capital is estimated to be US$124.7 million.

 

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Financial evaluations account for physical reclamation costs at various times in the LOM, for the development of a fund to address water treatment costs post reclamation and for special use securities associated with permanent access roads.

 

Working capital is estimated at one month of receivables and three months of payables and varies from year to year. The working capital will be recovered at the end of the mine life.

 

Pre-production construction period is estimated to be six years. NPV and IRR reported in this section are estimated at the start of this six-year period.

 

22.3.2Metal Price Scenarios

 

The 2022 PFS Base Case uses the three-year average metal prices as of June 2022 and a US$/Cdn$ exchange rate of 0.77. In addition to the 2022 PFS Base Case, two alternate cases are also presented: (i) a Recent Spot Case incorporating recent spot prices for gold, copper, silver, and the US$/Cdn$ exchange rate, and (ii) an Alternate Case that incorporates lower metal prices than used in the Base Case to demonstrate the 2022 PFS sensitivity to lower prices. The input parameters and pre-tax results of all scenarios are shown in Table 22.2.

 

Table 22.2Summary of the Pre-tax Economic Evaluations

 

  Unit 2022 PFS
Base
Case
2022 PFS
Recent Spot
Case
2022 PFS
Alternate
Case
Gold US$/oz 1,742.00 1,850.00 1,500.00
Copper US$/lb 3.53 4.25 3.00
Silver US$/oz 21.90 22.00 20.00
Molybdenum US$/lb 18.00 18.00 18.00
Exchange Rate US$:Cdn$ 0.77 0.77 0.77
Undiscounted NCF US$ million 38,636 46,070 27,854
NPV (at 3%) US$ million 20,210 24,357 14,210
NPV (at 5%) US$ million 13,454 16,403 9,194
NPV (at 8%) US$ million 7,420 9,294 4,717
IRR % 20.1 22.4 16.5
Payback years 3.4 3.1 4.1
Cash Cost/oz Au US$/oz 275 164 351
Total Cost/oz Au US$/oz 601 490 677
           

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22.4Post-tax Financial Evaluations

 

Seabridge engaged PwC in Toronto, Ontario to review the tax component of the model for the post-tax economic evaluation for this 2022 PFS with the inclusion of applicable income and mining taxes.

 

The following general tax regime was recognized as applicable in Q3 2022:

 

22.4.1Canadian Federal and BC Provincial Income Tax Regime

 

The federal and BC provincial corporate income taxes are calculated using the current enacted rates of 15% and 12% respectively. For both federal and provincial income tax purposes, capital and resource expenditures are accumulated in tax pools that can be deducted against mine income at different prescribed rates, depending on the type of capital expenditures.

 

All pre-production mine development expenses, Canadian resource property acquisition costs and the costs of mine shafts, main haulage ways, and other underground workings are considered Canadian development expense (CDE) and are accumulated in the CDE pool. The KSM Financial Model treats all such expenses as CDE.

 

Fixed assets acquired for the mine are accumulated in an undepreciated capital cost pool (Class 41) and are generally amortized at 25% on a declining balance basis.

 

CDE, except for costs with respect to an acquisition of a Canadian resource property, fixed assets, and Class 14.1 expenditures incurred after November 20, 2018 and before 2028 are eligible for an enhanced first-year allowance under the Accelerated Investment Incentive measure, which is factored into the KSM Financial Model.

 

22.4.2BC Mineral Tax Regime

 

The BC Mineral Tax regime is a two-part tax regime, with a 2% tax and a 13% tax.

 

The 2% tax is a minimum tax and is applied on “net current proceeds”, which is defined as gross revenue from the mine less mine operating expenditures. Hedging income and losses, royalties and financing costs are excluded from operating expenditures. The 2% tax is accumulated in a Cumulative Tax Credit Account (CTCA) and is fully creditable against the 13% tax.

 

All capital expenditures, both mine development costs and fixed asset purchases, and mine operating expenditures are accumulated in the Cumulative Expenditures Account (CEA), which is amortized at 100% and deductible for the purposes of the 13% tax.

 

The 13% tax is applicable on “net revenue”, which is defined as gross revenue from the mine less any accumulated CEA balance to the extent of the gross revenue from the mine for the year.

 

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A notional interest equal to 125% of the prevailing federal bank rate is calculated annually on any unused CEA and CTCA and is added to the respective balances.

 

BC Mineral Tax is deductible for federal and provincial income tax purposes.

 

22.4.3Taxes and Post-tax Financial Results

 

At the 2022 Base Case metal prices and exchange rate used for this study, total estimated taxes payable on KSM profits are US$14.7 billion over the 33-year LOM. The total estimated taxes payable by the mine in all scenarios provided are shown in Table 22.3.

 

The post-tax undiscounted annual cash flows are illustrated in Figure 22.1.

 

Figure 22.1Post-tax Undiscounted Annual and Cumulative Cash Flow

 

 

Source: Tetra Tech, 2022

 

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Table 22.3Component of the Various Taxes for all Scenarios

 

  Unit 2022 PFS
Base
Case
2022 PFS
Recent
Spot
Case
2022 PFS
Alternate
Case
Gold US$/oz 1,742.00 1,850.00 1,500.00
Copper US$/lb 3.53 4.25 3.00
Silver US$/oz 21.90 22.00 20.00
Molybdenum US$/lb 18.00 18.00 18.00
Exchange Rate US$:Cdn$ 0.77 0.77 0.77
Corporate Tax (Federal) US$ million 5,152 6,118 3,754
Corporate Tax (Provincial) US$ million 4,122 4,894 3,003
BC Mineral Tax US$ million 5,429 6,427 3,963
Total Taxes* US$ million 14,703 17,440 10,721

Note: *Totals may not add up due to rounding.

 

Post-tax financial results are summarized in Table 22.4.

 

Table 22.4Summary of Post-tax Financial Results

 

  Unit 2022 PFS
Base
Case
2022 PFS
Spot
Case
2022 PFS
Alternate
Case
Gold US$/oz 1,742.00 1,850.00 1,500.00
Copper US$/lb 3.53 4.25 3.00
Silver US$/oz 21.90 22.00 20.00
Molybdenum US$/lb 18.00 18.00 18.00
Exchange Rate US$:Cdn$ 0.77 0.77 0.77
Undiscounted NCF US$ million 23,933 28,630 17,133
NPV (at 3%) US$ million 12,264 14,889 8,467
NPV (at 5%) US$ million 7,944 9,814 5,238
NPV (at 8%) US$ million 4,061 5,254 2,332
IRR % 16.1 18.0 13.1
Payback years 3.7 3.4 4.3

 

Table 22.5 summarizes the 2022 PFS annual cash flow for the pre-production period, Years 1 to 7, and the LOM, providing mine and mill production, revenue projections, operating costs and capital costs, and undiscounted cash flows both before and after taxes.

 

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Table 22.5 2022 PFS Annual Cash Flow for Pre-production Period, Years 1 to 7 and LOM1,2

 

  Unit Pre-prod.
Period
Production Period
Years
-6 to -1
Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 LOM
Mine and Mill Production3
Waste Mined (incl. rehandle) Mt   216 211 87 12 82 78 45 2,417
Mill Feed Processed Mt   38 47 71 71 71 71 71 2,292
Grade
Gold (concentrate + doré) g/t   0.88 0.89 1.12 0.92 0.98 0.74 0.67 0.64
Copper (concentrate) %   0.23 0.22 0.20 0.21 0.21 0.23 0.16 0.14
Silver (concentrate + doré) g/t   2.5 2.7 3.6 3.3 3.6 2.8 2.4 2.2
Molybdenum (concentrate) ppm   44 49 49 49 43 54 69 76
Metal Recovered
Copper (concentrate) Mlbs   174.7 205.4 273.6 290.8 291.3 309.0 211.0 5,872.4
Gold (concentrate + doré) Moz   0.9 1.1 2.0 1.7 1.8 1.3 1.2 33.9
Silver (concentrate + doré) Moz   1.8 2.5 5.5 4.7 5.4 3.9 3.1 97.9
Molybdenum (concentrate) Mlbs   0.9 1.4 1.9 1.9 1.7 3.0 3.8 140.2
Total Revenue (NSR)
Copper Concentrate4 US$ million   1,718 2,090 3,443 3,045 3,207 2,652 2,077 57,359
Gold Doré and Silver Doré US$ million   266 352 919 700 764 552 577 18,838
Molybdenum Concentrate US$ million   14 23 30 30 26 46 59 2,188
Revenue (NSR) US$ million   1,998 2,464 4,392 3,775 3,997 3,251 2,713 78,385
Costs  
Total Royalties Payable US$ million (3) (27) (34) (87) (65) (81) (88) (73) (2,674)
NSR Net of Royalty US$ million   1,971 2,430 4,304 3,710 3,916 3,163 2,640 75,711
Total Operating Costs US$ million   (607) (688) (899) (692) (822) (778) (782) (26,035)
Operating Cash Flow US$ million   1,364 1,742 3,405 3,018 3,094 2,385 1,858 49,676
Total Capital Costs US$ million (6,432) (723) (300) (162) (109) (68) 25 27 (9,642)
PST US$ million (83) (9) (4) (1) (1) (1) (0) (0) (125)
                  table continues…

 

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  Unit Pre-prod.
Period
Production Period
Years
-6 to -1
Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 LOM
Reclamation- water treatment fund US$ million (9) (70) 0 0 0 0 (93) (23) (1,273)
Pre-tax Cash Flow US$ million   562 1,439 3,242 2,907 3,025 2,317 1,862 38,636
BC Mining Tax US$ million   (28) (36) (70) (62) (368) (318) (249) (5,429)
BC Income Tax US$ million   - - (49) (297) (282) (213) (167) (4,122)
Federal Income Tax US$ million   - - (62) (371) (353) (267) (208) (5,152)
Total Taxes US$ million   (28) (36) (181) (730) (1,002) (798) (624) (14,703)
Net Cash Flow US$ million   535 1,403 3,061 2,178 2,022 1,519 1,238 23,933

Notes:1. The complete annual cashflow schedule is presented in Figure 22.1.

2. The complete annual mine production schedule is presented in Table 16.9 and Figure 16.12.

3. The metal grades presented herein are the grades of the mill feed that will eventually report to the concentrate or doré.

4. Includes gold and silver in concentrate form.

5. Royalties include IBA payments.

6. Sums may not add due to rounding.

 

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22.5Sensitivity Analysis

 

Tetra Tech investigated the sensitivity of NPV and IRR to the key variables. Using the 2022 PFS Base Case as a reference, each of key variables was changed between -30% and +30% in 10% increments while holding the other variables constant.

 

Sensitivity analyses were carried out on the following key variables:

 

·gold, copper, silver, and molybdenum metal prices

 

·exchange rate

 

·capital costs

 

·operating costs.

 

The analyses are presented graphically as financial outcomes in terms of post-tax NPV, and IRR. The NPV is most sensitive to gold price and exchange rate, followed by operating costs, copper price, and capital costs. The IRR is most sensitive to exchange rate, capital costs, and gold price, followed by copper price and operating costs. In general, sensitivity to metal price is roughly equivalent to sensitivity to metal grade. Financial outcomes are relatively insensitive to silver and molybdenum prices. The NPV and IRR sensitivities are presented in Figure 22.2 and Figure 22.3, respectively.

 

Figure 22.2Sensitivity Analysis of Post-tax NPV at a 5% Discount Rate

 

Source: Tetra Tech, 2022

 

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Figure 22.3Sensitivity Analysis of Post-tax IRR

 

Source: Tetra Tech, 2022

 

22.6Royalties

 

The following royalties are included in the economic analysis.

 

1.1% of the NSR payable to Newmont, capped at US$3.5 million, with a predetermined buyout option. The full amount of the buyout option is paid in Year 1 in the financial model. Refer to Section 4.2 for details.

 

2.1.5% of the NSR payable to Pretium (East Mitchell only)

 

3.60% of the gross silver royalty payable to Sprott

 

4.Commitments in Impact Benefit Agreements listed in Section 20.1.5

 

22.7Smelter Terms

 

The copper concentrate smelter terms and molybdenum concentrate smelter charges that have been applied in the economic analysis are presented in Section 19.1 and 19.2 of this report, respectively.

 

Gold and silver doré will generally include payment terms as follows:

 

·Gold: pay 99.8% of content less a refining charge of US$1.00/accountable oz.

 

·Silver: pay 90.0% of content less a refining charge of US$1.00/accountable oz.

 

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22.8Miscellaneous Costs and Charges

 

An assumption is being made that the customer base will be in Asia. The copper concentrate will be shipped in bulk via port facilities in Stewart, BC, and the molybdenum concentrate will be packed in concentrate bags and shipped inside sea containers via port facilities in Prince Rupert, BC. Transportation costs for the copper and molybdenum concentrate are listed below:

 

·Copper concentrate:

 

̶trucking: US$35.28/wmt

 

̶port storage and handling: US$16.94/wmt

 

̶ocean transport to Asian port: US$63.50/wmt

 

̶moisture content: 9%.

 

·Molybdenum concentrate:

 

̶trucking: US$81.15/wmt

 

̶port storage and handling: US$29.63/wmt

 

̶ocean transport to Asian port: US$87.68/wmt

 

̶moisture content: 5%.

 

Gold and silver doré transportation cost is assumed to be US$1.00/oz.

 

An insurance rate of 0.125% was applied to the provisional invoice value of the concentrates and doré to cover land-based and ocean transport from the mine site to the smelter.

 

A US$8.50/dmt charge was applied for marketing and services provided by the Owner’s representative. Duties would include attendance during vessel unloading at the smelter port, supervising the taking of samples for assaying, and determining moisture content.

 

An overall weight loss of 0.1% was applied in the economic analysis.

 

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23.0Adjacent Properties

 

There are no relevant adjacent properties to the KSM Property that is the subject of this Report.

 

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24.0Other Relevant Data and Information

 

The 2022 PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the results of the 2022 PEA will be realized. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

 

The 2022 PEA is a stand-alone mine plan that has been undertaken to evaluate a potential future expansion of the KSM mine to the Iron Cap and Kerr deposits after the 2022 PFS mine plan has been completed. None of the Mineral Resources incorporated into the 2022 PEA mine plan have been used in the 2022 PFS mine plan.

 

The 2022 PEA is a scoping level of study based on the Mineral Resources stated in the Kerr and Iron Cap Mineral Resource estimate in Section 14.

 

The 2022 PEA is primarily an underground block cave mining operation supplemented with a small open pit and is planned to operate for 39 years with a peak mill feed production of 170,000 t/d. The 2022 PEA mine site general arrangement is shown in Figure 24.1. The numbering scheme for the sub-sections of this part of the Report follow Items 16 through 22 of Form 43-101F1.

 

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Figure 24.1Mine Site General Arrangement (2022 PEA)

 

Source: Tetra Tech (2022)

 

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24.16Mining Methods

 

24.16.1Net Smelter Return Block Model

 

The NSR per tonne for each block in the block model used for the PEA mine planning is determined using the Mineral Resources discussed in Section 14.0. It is net of offsite concentrate treatment and smelter charges and inclusive of onsite mill recovery. It is used as a cut-off item for break-even mill-feed/waste selection, as well as for grade bins used to optimize cash flow in the open pit production scheduling.

 

NSR is estimated using Net Smelter Price (NSP) and process recoveries. Metal Prices and exchange rate in Table 24.1 and typical smelter terms, off site costs and royalties are used to estimate the NSP. Process recoveries have been determined using the metallurgical projections described in Section 13, with the exclusion of CIL gold recovery from mill feed derived from Iron Cap and Kerr zones.

 

Table 24.1Metal Prices for PEA NSR Calculation

 

  Metal Price
(US$)
Cu 2.70/lb
Au 1,200/oz
Ag 17.50/oz
Mo 9.70/lb
Exchange Rate (US$:Cdn$) 0.83
   

24.16.2Open Pit Mining Method

 

The open pit development is designed as a conventional truck-shovel operation with 360 t trucks and 56 m3 and 40 m3 shovels. Mill feed will be hauled to the Mitchell primary crusher and waste rock will be hauled to the Mitchell open pit backfill.

 

Pit Design

 

Kerr open pit has been designed to supplement block cave mill feed during the ramp up of the block cave production. The pit limit has been selected by using the first phase of the Kerr pit design from previous studies. The Kerr pit design incorporate open pit slope design parameters based on geotechnical site investigations, available local and regional geological data, and well-established geotechnical design methods.

 

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The mine design for the Kerr pit phase is shown in Figure 24.2.

 

Figure 24.2Kerr Pit Design – Plan View

 

 

Source: MMTS (2022)

 

Sub-set of Mineral Resources Within PEA Pit

 

The Mineral Resources summarized in Table 24.2 is a subset of the Mineral Resources and quantities that are included in Section 14. The Mineral Resources within the PEA pit uses an NSR cut-off grade of Cdn$10.75/t.

 

Table 24.2Mineral Resources Within the PEA Open Pit Mine Plan

 

Pit

Mill Feed

(Mt)

Au

(g/t)

Cu

(%)

Ag

(g/t)

Mo

(ppm)

Indicated 117 0.26 0.51 1.4 1.8
Inferred 7 0.74 0.09 1.5 0.6

 

Waste Rock Facilities

 

All open pit waste is placed in the Mitchell mined out pit as backfill.

 

Production Scheduling (Open Pit)

 

There is one year of pre-stripping before production commences. Open pit production will be carried from PEA Years 1 to 5 using an already established KSM open pit mining fleet to supplement Iron Cap mill feed ramp up. Open pit production is completed before Kerr underground production begins.

 

Open pit scheduling has been carried out with the MineSight MPSO with the primary program objective to maximize the utilization of the available open pit mine fleet.

 

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Kerr Open Pit Mine Operations

 

KSM Kerr open pit mining operations will employ bulk mining methods and large capacity equipment that are suitable for the rates of mining that are proposed in the PEA mine plan.

 

24.16.3Iron Cap Mining Methods

 

Mining Method Selection

 

Based on mining costs, deposit grade, geometry, and depth, block caving was selected as the preferred mining method. Iron Cap has been designed to operate battery-electric loaders and electric trains for material handling and employs both tele-operation and automation for these units. This enables a reduction in the number of primary ventilation intake and exhaust drifts due to less ventilation demand and cost savings from lower diesel fuel, labour and equipment maintenance costs.

 

Geotechnical and Caving Geomechanics Considerations

 

The characterization of the Iron Cap rock mass focuses on the rock in and around the extraction and undercut levels of the proposed block cave mine (870 m elevation) and on the mineralized rock above this that will be caved.

 

Characterization of the rock was based on core photographs and data collected for exploration drill holes, detailed geotechnical data collected for drilling programs carried out by BGC in 2010 (BGC, 2011), and an interpreted geological model provided to WSP Golder by Seabridge.

 

Three geotechnical holes were drilled in the Iron Cap deposit in 2010 and no additional geotechnical holes have been drilled since then. A qualitative geotechnical assessment of the core from a number of exploration holes drilled in 2017 was undertaken by WSP Golder personnel in September 2017. This indicated that the geotechnical quality of the rock mass was good to very good, and the quality was very uniform in the area that is being proposed to be mined by block caving. The overall conclusion was that the geotechnical assessment undertaken by WSP Golder for previous studies (Golder, 2012) remains valid for this 2022 PEA.

 

The caveability assessments made using Laubscher’s and Mathews’ methods for prior studies indicated that the size of the footprint required to initiate and propagate caving is less than approximately 200 m. This minimum footprint size is significantly smaller than the size of the footprint of the deposit that can potentially be mined economically by caving. This fact, together with the generally large-size continuous nature of the deposit indicates that the Iron Cap deposit is amenable to block caving.

 

The cave mining will draw down the mineralized rock and a significant depression will develop on surface above the production footprint in the form of a crater. The crater typically develops on surface above and slightly laterally beyond the footprint of the production horizon of the cave mining. The top section of the crater is a relatively steep escarpment (60° to 70°) that is marginally stable but comprised of nominally in place dilated rock. Deeper within the crater is failed broken rock that has progressively sloughed from the rim of the crater. This rock rills down to the bottom of the crater at angle of repose.

 

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The Iron Cap cave will undercut parts of the overlying ice sheet. The ice will break up as part of the surface disturbance created by the cave and progressively melt as it is drawn down. The total volume of water that enters the cave from the melting ice is relatively small. Further studies will be required in the future to assess the need to drain free water that may be present immediately beneath the ice sheet prior to the start of caving.

 

Cave Modelling and Production Scenarios

 

GEOVIA’s PCBC Footprint Finder software (FF) was used to assess a range of footprint shapes and elevations for a range of NSR shut-off values in excess of the preliminary operating cost estimate of Cdn$11.50 per tonne. This cost is comprised of Cdn$3.80/t mining cost, and Cdn$7.70/t processing and general and administrative (G&A) costs. These preliminary costs represent the initial estimates that were made to undertake the FF assessments. The mining operating cost was subsequently refined to Cdn$5.64/t (US$4.34/t). As well the process and G&A operating cost was subsequently refined to Cdn$9.08/t.

 

The NSR values were obtained from the block model discussed in Section 24.16.1. The primary parameters used in the FF assessments are presented in Table 24.3. A maximum column height (height of draw, HOD) of 750 m was used in the FF assessments.

 

Table 24.3Iron Cap Footprint Finder Input Parameters

 

Footprint Finder Parameter Value Comments
Discount Rate 5% Provided by Seabridge
Maximum Height of Draw 750 m Industry experience
Minimum Height of Draw 195 m From caving geomechanics assessments

Vertical Incremental

Capital Cost

Cdn$180,000 per m WSP Golder benchmark from previous studies

Lateral Incremental

Capital Cost

Cdn$2,100 per m2 WSP Golder benchmark from previous studies

Preliminary Operating Cost

(Mine, Mill and G&A)

Cdn$11.50 per tonne $3.80 for mining (WSP Golder) and $7.70 for processing and general and administration (Seabridge)
Maximum Draw Rate 80 m per year From industry experience
Drawpoint Opening Rate 120 per year Construction of 10 drawpoints per month – industry experience
Production Ramp up Rate Avg. 5 Mt per year Industry experience (3 to 6 Mt per year)
Maximum Production Rate 90,000 tpd Achievable rate based on the 80 m/yr. draw rate, 120 drawpoints constructed per year and available footprint area

 

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The shut-offs used in the FF assessments ranged from Cdn$16/t to Cdn$24/t which represent a premium over the preliminary operating cost estimate. From this a footprint at elevation 870 m was selected based on an NSR shut-off of Cdn$20/t. Part of the selection of the footprint elevation was to provide for gravity drainage of inflow water from direct precipitation, surface runoff, and operational water. This design generated sufficient tonnes to feed the mill in combination with other mill feed sources over a sustained period and at a production rate of 90 kt/d. The total production is 743 million tonnes. A second lower cave lift was assessed but the footprint was not sufficiently large to warrant further study. Figure 24.3 shows the final footprint profile at 870 m elevation. The warmer colors shown in this figure represent higher column values. The production schedule for the Cdn$20/t NSR shut-off is presented later in Section 24.16.3.

 

Figure 24.3Iron Cap Footprint for Lift at 870 m Elevation

 

Note: (WSP Golder 2022) Grid system is based on UTM coordinates. Warmer colours indicate higher column NSR values and cooler colours indicate lower column NSR values.

 

Design Criteria and Layout

 

The block cave design for the Iron Cap footprint is at the 870 m elevation. The shape of footprint is an irregular oval, with an area of 570,000 m2, a length of approximately 1 km, and a width at the widest section of approximately 700 m

 

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The heights of draw through which the production tonnes are extracted average approximately 500 m ranging from 250 m to 750 m, where the lower values reflect the western side of the footprint that is developed earlier in the mine life and the higher values reflect the eastern side of the footprint that is developed late in the mine life. Most of this geometry is dictated by the position of the Sulphurets Thrust Fault which is the structure which constrains economic mineralization below it.

 

The development direction for the Iron Cap cave starts from the middle of the eastern lobe and propagates away from this middle to the footprint edge; and also generally from west to east, such that draw bells on the western half of the footprint are developed first and draw mill feed columns whose heights are generally lower than the average height across the footprint.

 

Personnel, material, and supplies will access the underground through the primary access drift adjacent to the Mitchell OPC. For this design, the MTT has been slightly re-aligned towards the south adjacent to the Iron Cap deposit to avoid intersecting with the very weak STF and the adjacent mylonitic zone. Immediately adjacent to the MTT is a rail spur to allow the haulage trains to be loaded under two surge bins that are fed by the Iron Cap conveyor. The conveyor is fed by two gyratory crushers located just outside the northern end of the footprint.

 

The haulage level maintenance shops, underground offices and warehouse facilities are located at the north end of the footprint.

 

Two fresh air portals and two exhaust portals are planned on the north slope of the Mitchell Valley as seen in Figure 24.4. A third fresh air intake adit will be collared in the area of the Mitchell OPC to the southwest of the footprint. It is required to provide a flow-through ventilation circuit during the pre-production development phase before the other drifts can be excavated south from the footprint to breakthrough on the north slope. Developing the tunnels northward from the north valley slope into the mountain was considered unsafe due to the high geohazard and avalanche risks on this slope. The third adit will remain an active intake and access adit throughout the mine life.

 

The fresh air adits will connect to perimeter drifts constructed around the entire mine footprint to provide fresh air to the mine workings. Exhaust raises on the extraction level will connect down to an exhaust level that then connects to the two adits that exhaust air to the north slope of the valley. The exhaust portals are downwind of the fresh air portals to avoid intake of air contamination and are separated approximately 340 m from one another.

 

The Iron Cap design has multiple drifts that can act as emergency egress. The primary emergency egress will be the train tunnel or fresh air adit of the MTT, whichever is accessible. If both are inaccessible, it will be possible to exit the mine through one of the fresh air drifts that connect to the north slope of the Mitchell Valley.

 

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Figure 24.4Plan View of Iron Cap Mine with Contour

 

 

Source: WSP Golder (2022)

 

Figure 24.5 shows the typical level arrangement for the Iron Cap mine. The mine development occurs on the following seven levels:

 

the pre-conditioning level is planned to provide access for in situ fracturing of the rock mass prior to caving. From the grid of drifts on this level, a series of boreholes will be drilled, and hydraulic fracturing will be used to generate fractures in the rock mass that it is proposed to cave

 

the undercut level is comprised of a series of parallel drifts to perform drilling and blasting required to develop void and initiate caving. The level is located

40 m below the pre-conditioning level

 

the extraction level is comprised of parallel drifts spaced 18 m apart where the drawpoints and dumping points are located. Battery-electric LHD equipment will extract the caved material from the drawpoints and haul it to the nearest ore pass dump point. This extraction level at Elevation 870 m is 20 m below the undercut level

 

the fresh air level is 20 m below the extraction level and conveys fresh air to the other levels via vertical ventilation raises

 

the exhaust ventilation level is 5 m below the fresh air level and this level exhausts air by vertical ventilation raises from the extraction level

 

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the haulage level is 25 m below the exhaust level. It receives the material dumped by LHD’s into ore passes on the extraction level. Each ore pass from the extraction level is fitted with a stationary rock breaker to size the material that gravity feeds down to chutes that load rail cars. Electric locomotives haul the loaded rail cars to one of two gyratory crushers located just outside the northern end of the footprint

 

the water drainage level is located 25 m below the haulage level and it collects drainage and inflow water and directs it to the 1,314 m long Dewatering Tunnel that gravity drains to the North Pit Wall Dewatering Adit (NPWDA).

 

Figure 24.5Typical Level Arrangement for Iron Cap

 

Source: WSP Golder (2022)

 

An El Teniente-type drawpoint layout is planned on the extraction level. The routing for the LHD haulage to the ore passes is shown in Figure 24.6. The design incorporates extraction drifts that are 5.0 m wide by 4.6 m high and spaced 30 m apart, with drawbell drifts that are similarly sized, aligned at a 60° angle to the extraction drifts, spaced 18 m apart. The drawbells are created by drilling and blasting and then two drawpoints per drawbell are constructed. These drawpoints are supported with heavy support in the brows in order to withstand the attrition and erosion of the mineralized material as it is mucked. The resulting drawpoint spacing is 18 m by 15 m which is appropriate for the expected fragmentation.

 

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Figure 24.6Drawpoint Configuration for Iron Cap

 

 

Source: WSP Golder (2022)

 

Overall, the mine plan is based on there being 7 different levels and approximately 200 km of drifts and raises. Table 24.4 summarizes the various types of drift development and the sizes of these excavations.

 

Table 24.4Main Drift Types and Dimensions

 

Type Dimensions
Drawpoint and Extraction Drifts 5.0 m W x 4.6 m H
Perimeter, internal ramps 6.0 m W x 5.5 m H
Pre-conditioning, Undercut, Dewatering Drifts 5.0 m W x 4.6 m H
Access Tunnels, Conveyor Drifts 6.0 m W x 6.0 m H
Haulage Level 6.0 m W x 6.0 m H
Main Fresh/Return Air Drifts 7.5 m W x 7.5 m H

 

Vertical development will be constructed primarily using raisebore machines and raisebore-and-slash techniques.

 

Table 24.5 summarizes the various types of vertical development and the sizes of these excavations.

 

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Table 24.5Vertical Development Dimensions

 

Type Diameter (m) Type
Ore Passes and Haulage Level Ventilation Raises 4 Bored Raise
Footprint, Undercut, and Pre-conditioning Level Ventilation Raises 2.5 Bored Raise
Ore Bins (including MTT ore bin) 10 Bored Raise and Slashed

 

Ventilation

 

The quantity of required airflow was determined using a VentSim model of the mine design and was based on the requirement for 0.06 m3/s of ventilating air for each kilowatt of power of diesel-powered equipment and 0.025 m3/s per kilowatt of electrically powered equipment. Mine equipment and engine utilization factors were also considered in determining total ventilation needs.

 

The peak airflow requirement was estimated to be 1,075 m3/s, which is sufficient to dilute all noxious gases, particulate matter, and heat produced by the mining equipment and the operational activities on each mining level. This equates to a rate of 0.011 m3/s per tonne of material mined per day and is approximately one-half of a typical block cave mine using only diesel-powered equipment. This reduced ventilation demand results in fewer primary intake and exhaust drifts and less main fan capacity and power demand.

 

There are three main intakes and two main exhausts that are 7.5 m by 7.5 m in cross-section. There are 3 axial fans at each of 3 intake drifts with a peak power demand of 1,865 kW. Raises that are 2.5 m in diameter connect the extraction level to the exhaust drifts.

 

Heating of mine air in the winter months is included in the design and cost estimates, and this will be accomplished using propane mine heaters located at each of the two main fan installations. A mine air heating system of 15 MW capacity is required to heat 1,075 m3/s of air from an average minimum temperature of -5.7°C to 2°C for an average of six months per year (November through April). The average propane requirement is 3.5 million liters per year during production. Figure 24.7 is a schematic of the ventilation system showing major airflow directions.

 

Figure 24.7Ventilation Schematic for Iron Cap Mine (Section Looking West)

 

Source: WSP Golder (2022)

 

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Mine Dewatering

 

The maximum mine water inflow estimate for Iron Cap is 6 m3/s. This is based on factoring the more detailed assessment that was undertaken from first principles. The inflow water includes groundwater and surface water inflows (including ice melt from ice that is drawn down with the caved rock when the icefield is undercut and then melts), and water that is introduced to the mine for operations. The inflow from melting ice will have a relatively insignificant impact on total water inflows. The inflows to Iron Cap are conveyed by gravity via a system of sumps and drainholes from the various levels to a 1.3 km drift located on the southwest corner of the footprint that connects to the North Pit Wall Dewatering Adit (NPWDA). The NPWDA flows into the Mitchell Valley Depressurization Tunnel (MVDT) which then flows to the WSF.

 

All drifts have been graded so that water flows towards collection sumps which will feed into the Dewatering Tunnel. Sump pumps will be used locally as required. The water management system has a nominal capacity of 6 m3/s (95,000 gpm).

 

2022 PEA Production Schedule

 

The proposed production schedule as developed using FF software is shown in Figure 24.8. Note that the years shown apply to the combined open pit and underground multi-mine plan of the KSM 2022 PEA mine production. Iron Cap pre-production construction will take 4 years with the first cave production being in PEA Year 1.

 

A production ramp-up period of 6 years is governed by the draw point opening rate of 120 per year and maximum draw rate of 80 m per year (220 mm per day). Both are consistent with industry experience. The sustained peak production rate estimated using FF is 90 kt/d. There is also a 6-year ramp-down period as drawpoints are depleted.

 

Figure 24.8Iron Cap Mine Production Plan

 

 

Source: WSP Golder (2022)

 

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Only Indicated and Inferred Mineral Resources are included in the production plan. The production tonnes and grades are comprised of mineralized material included in the Mineral Resources reported in Section 14, mineralized dilution taken at grade (the portion of Indicated and Inferred material that is below the operating costs cut-off) and includes small percentage of material at zero grade. All unclassified materials are treated as waste by setting the grades to zero. Table 24.6 shows the distribution of production tonnes and grades in the 2022 PEA mine plan.

 

Table 24.6Iron Cap Production Tonnes and Grade in 2022 PEA Mine Plan

 

  Measured Indicated Inferred
Tonnes (millions) - 57.7 685.0
NSR (Cdn$/t) - 32.83 36.70
Au (g/t) - 0.62 0.58
Cu (%) - 0.28 0.36
Ag (g/t) - 3.2 3.0

Note: Includes mineralized dilution (the portion of Indicated and Inferred Mineral Resources that is below the operating costs) and includes small percentage of material at zero grade.

 

The mining sequence supporting the production plan is shown in Figure 24.9. The cave is initiated in the region of the footprint that has the highest NSR values to maximize early value. The mining then propagates radially from this point and continues down the southeastern limb of the footprint.

 

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Figure 24.9Proposed Drawpoint Development Sequence for the Iron Cap Mine

 

 

Note: Legend is drawpoint number Source: WSP Golder (2022)

 

Development Schedule

 

The pre-production development phase of the Iron Cap mine is estimated to be 4 years, which includes mine access, some initial footprint development and construction of major mine infrastructure, such as the underground crusher, material handling system, shops, dewatering system, and primary ventilation. The development schedule was generated using Deswik software based on an advance rate of 4 m/day and a maximum of three active headings in the first two years, six active headings in PEA Year 3, and nine active headings in PEA Year 4.

 

The overall development strategy is to provide access to the crusher location as soon as possible while prioritizing the installation of the materials handling system, haulage system and shop areas. Following this, extraction drifts, drawpoints, and undercut and pre-conditioning levels will be advanced to allow for cave initiation. Several ventilation airways will also be advanced southward towards the Mitchell Valley to establish the flow-through ventilation circuit in advance of production.

 

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Blasting and Explosives

 

It is assumed that explosives will be stored in a surface magazine in the Mitchell Valley to service the Iron Cap mine. The explosives will be delivered to the underground mine via the access tunnel as required. Small underground magazines will be established to provide daily supplies of explosives, and explosives will be distributed on a shift basis for five main purposes:

 

preparation for the development of mineralized areas

 

draw bell construction

 

undercutting

 

development in barren material

 

secondary blasting of boulders and hang ups.

 

Mining Equipment

 

The peak primary and support equipment requirements for the proposed Iron Cap mine are summarized in Table 24.7.

 

Table 24.7Peak Mobile Equipment Requirements for Iron Cap

 

Development

Equipment

Unit

Production

Equipment

Unit

Support

Equipment

Unit
Jumbo 12 LHD (BEV) 24 Grader 2
Haul Truck (BEV) 8 Drill Rig 4 Lube Truck 3
LHD (BEV) 5 Water Cannon 4 Electricians Truck 3
Bolter 16 Block Holer 8 Telehandler 4
Explosive Loader 5 Mobile Rockbreaker 5 Mobile Crane 2
Scissor Lift 7 Boom Truck 1 Mine Ambulance 1
Shotcrete Sprayer 4 Locomotive (Electric) 5 Fire Truck 1
Transmixer 5 Shunting Locomotive (Electric) 1 Mine Rescue Truck 1
Raisebore 3 Rail Car 78 Fan/Pipe Handler 1
        Backhoe 3
        Water Truck 1
        Sludge Truck 1
        Cable Bolter 4
        Big Personnel Carrier 6
        Small Personnel Carrier 16

Note: BEV is battery-electric vehicle.

 

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Mine Workforce

 

The estimated workforce for Iron Cap is 658 persons at the peak of the pre-production construction period with an average of 600 persons during the production period.

 

24.16.4Kerr Mining Methods

 

Mining Method Selection

 

Based on mining costs, deposit grade, geometry and depth, block caving was selected as the preferred mining method for the lower part of the Kerr deposit. Kerr has been designed for using diesel powered loaders and haulage trucks for material handling.

 

Geotechnical and Caving Geomechanics Considerations

 

Dedicated geotechnical drilling or studies have not yet been conducted for the Kerr resource area. Current geological exploration drilling indicates similar rock conditions to the geotechnical information that were gathered from the Mitchell deposit, and for the purposes of the 2022 PEA it is assumed the rock characteristics will be similar.

 

The caveability assessments made using Laubscher’s and Mathews’ methods indicate that the size (diameter) of the footprint required to initiate and propagate caving is less than 200 m. This minimum footprint size is significantly smaller than the size of the footprint of the deposit that can potentially be mined economically by caving. This fact, together with the general large-sized, continuous nature of the deposit, suggests that the lower Kerr deposit is amenable to cave mining.

 

The cave mining will draw down the mineralized rock and a significant depression will develop on surface above the production footprint in the form of a crater. The crater typically develops on surface above and slightly laterally beyond the footprint of the production horizon of the cave mining. The top section of the crater is a relatively steep escarpment (60° to 70°) that is marginally stable but comprised of nominally in place dilated rock. Deeper within the crater is failed broken rock that has progressively sloughed from the rim of the crater. This rock rills down to the bottom of the crater at angle of repose.

 

Life of mine underground infrastructure such as conveyors, vent drifts, ramps, and vent shafts are not planned within or near any areas potentially disturbed by caving.

 

Cave Modelling and Production Scenarios

 

GEOVIA’s PCBC Footprint Finder software (FF) was used to assess a range of footprint shapes and elevations for a range of NSR shut-off values in excess of the preliminary operating cost estimate of Cdn$13.60 per tonne. This cost is comprised of Cdn$6.00/t mining cost, and Cdn$7.60/t processing and general and administrative (G&A) costs. These preliminary costs represent the initial estimates that were made to undertake the FF assessments. The mining operating cost was subsequently refined to Cdn$6.82 (US$5.25). As well the process and G&A operating cost was subsequently refined to Cdn$9.08/t.

 

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The NSR values were obtained from the block model discussed in Section 24.16.1. The primary parameters used in the FF assessments are presented in Table 24.3. A maximum column height (height of draw, HOD) of 750 m was used to identify and locate the footprint for the upper lift, and 500 m was used for a subsequent lower lift. The proposed Kerr Pit was also included in the assessment. The primary FF input parameters are presented in Table 24.8. The assessment was based on a combined 2-lift operation.

 

Table 24.8Kerr Footprint Finder Input Parameters

 

Footprint Finder Parameter Value Comments
Discount Rate 5% Provided by Seabridge
Maximum Height of Draw 750 m 750 m used to define footprint and for production schedule, based on industry experience
Minimum Height of Draw 195 m From experience
Minimum Allowable Footprint Width 200 m Based on geotechnical/caveability assessments
Vertical Incremental Capital Cost Cdn$180,000 per m WSP Golder benchmark from previous studies
Lateral Incremental Capital Cost Cdn$2,100 per m2 WSP Golder benchmark from previous studies

Preliminary Operating Cost

(Mine, Mill and G&A)

Cdn$13.60 per t Cdn$6.00 for mining (WSP Golder) and Cdn$7.60 for processing and general and administration (Seabridge)
Maximum Draw Rate 80 m per year From industry experience
Drawpoint Opening Rate 120 per year Construction of 10 drawpoints per month – industry experience
Production Ramp up Rate (PRC) Avg. 5 Mt per year Industry experience (3 to 6 Mt per year)
Maximum Production Rate 80,000 tpd Achievable rate based on the 80 m per year draw rate, 120 drawpoints constructed per year and available footprint area

 

A range of NSR shut-offs values that included premiums over the preliminary operating cost estimate was assessed using FF. From this, footprint elevations of 625 m and 130 m were selected based on an NSR shut-off of Cdn$18/t. This generated sufficient tonnes to feed the mill in combination with other mill feed sources over a sustained period and at a production rate of 80 kt/d. The production from Lift 1 is 525 million tonnes and the production from Lift 2 is 300 million tonnes. The footprints for these two levels are shown in Figure 24.10 and Figure 24.11. The warmer colours shown in these figures represent higher column values. The final footprint boundaries maintain a minimum footprint width of 200 m to initiate and propagate caving (grid spacing on plan is 200 m).

 

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Figure 24.10Plan View of Kerr 625 m Lift 1 Footprint.

 

 

WSP Golder (2022) Grid system is based on UTM coordinates. Warmer colours indicate higher column NSR values and cooler colours indicate lower column NSR values.

 

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Figure 24.11Plan View of Kerr 130 m Lift 2 Footprint.

 

 

WSP Golder (2022) Grid system is based on UTM coordinates. Warmer colours indicate higher column NSR values and cooler colours indicate lower column NSR values.

 

Design Criteria and Layout

 

The block cave mine design is based on two separate Lifts with a combined area of 775,000 m2. Lift 1 is at the 625 m elevation and Lift 2 is at the 130 m elevation. Lift 1 tonnes are based on a starting topography that has the Kerr pit exploited.

 

Lifts 1 and 2 are both accessed through a series of shared portals and ventilation adits along the northwestern base of the mountain, facing Sulphurets Creek. All Lifts connect to a 7 km common inclined UG conveyor system which passes under the Sulphurets Creek and connects directly to an ore bin above the MTT material handling system which supports all mines at the KSM site. Mining on each lift assumes panel caving with a post undercutting method, with load-haul-dump (LHD) equipment for extraction, and a material handling system of muck passes, chutes, truck haulage, two crushers and conveyors.

 

The ventilation system design assumes five main intake adits and four main exhaust adits. Lift 2 is connected to the common ventilation adit system via bored shafts and decline ramps. Two of the main intake adits also serve as the primary access for the mine and are connected to the decline ramp system for the subsequent lower lift.

 

Surface camps, shops, warehouses, offices, electrical substation, water distribution, and other surface infrastructures are located just outside the portal area just below 600 m elevation. These facilities will be used in conjunction with similar facilities established for the overall KSM area to support the other mining operations, i.e., Mitchell, Iron Cap, Sulphurets, etc. Figure 24.12 shows a plan view of the mine layout projected to surface.

 

The Kerr design has multiple drifts that can act as an emergency egress. The primary emergency egress will be one of the fresh air adits. If these adits are inaccessible, it will be possible to exit the mine through the service adit. Egress from Lift 2 will be up the ramp or via the service ramp parallel to the conveyor decline.

 

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Figure 24.12Kerr Mine Plan View

 

 

Note: Spacing between Lifts 1 and 2 is 495 m. The distance from the portals to the footprint of Lift 1 is approximately 2 km. Source: WSP Golder (2022)

 

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Lift 1 has an area of 446,000 m2. The footprint shape is an irregular oval shape spanning 1,290 m in length with variable widths of up to 500 m, tapering at both ends. The extraction level is at elevation 625 m, which is slightly higher than the portal elevation of 600 m and thus does not require major ramps or shafts for access or ventilation. Mine access is primarily through two of the intake portals, one of which directly connects to the footprint area as a main ventilation intake drive, and the other which eventually parallels the UG conveyor as a service decline that connects to the crusher area. The total production tonnage from lift 1 will be 525 million tonnes. Figure 24.13 shows the general arrangement of Lift 1.

 

Figure 24.13Kerr Lift 1 (625L Mine Area) General Arrangement

 

 

Source: WSP Golder (2022)

 

Lift 2 is smaller in tonnage and footprint size than Lift 1 with an area of 329,000 m2. The footprint shape is an irregular and approximately 1,080 m in length with widths varying from 200 m up to 700 m, tapering at both ends. The extraction level is at elevation 130 m. It requires two decline ramps for access and muck conveyance, and seven bored shafts for ventilation. The total production tonnage from Lift 2 will be 300 million tonnes. Figure 24.14 shows the general arrangement of Lift 2.

 

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Figure 24.14Kerr Lift 2 (130L Mine Area) General Arrangement

 

 

Source: WSP Golder (2022)

 

Figure 24.15 shows the typical level arrangement for the Kerr underground mine.

 

Figure 24.15Typical Level Arrangement for Kerr Underground

 

Note: Light blue parallel undercut drifts and red parallel extraction panel drifts are spaced 30 m apart. Dark blue drawpoints are spaced 18 m apart. Source: WSP Golder (2022)

 

The underground mine development occurs on the following seven levels for each lift:

 

the pre-conditioning level is comprised of parallel drifts from which hydrofracking is performed to improve fragmentation. It is accessed by one perimeter drift on the west side only

 

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the undercut level is comprised of a series of parallel drifts to perform drilling and blasting required to develop a void and initiate caving. The undercut level is 40 m below the preconditioning level

 

the extraction level is comprised of parallel drifts spaced 18 m apart where the draw points and dumping points are located. Load-Haul-Dump (LHD) equipment extract the caved material from the draw points and haul it to the nearest dumping points for the ore pass. The extraction level is 17 m below the undercutting level

 

the exhaust ventilation level is 22 m below the extraction level and this level exhausts air from the other levels via vertical ventilation raises from the extraction level

 

the intake, or fresh air, ventilation level is 16 m below the exhaust level and conveys fresh air to the other levels via vertical ventilation raises

 

the haulage level is 17 m below the fresh air level. It receives the material dumped by diesel powered LHD’s into ore passes on the extraction level. Each ore pass is fitted with a stationary rockbreaker to size dumped material and a loading chute at the bottom to enable truck loading. Diesel haul trucks transport the material to one of two gyratory crushers located just outside the footprint

 

the dewatering level collects all drainage water via gravity through a series of sumps and boreholes connecting all the levels. The dewatering level is 16 m below the haulage level.

 

On the extraction level, an El Teniente-type drawpoint layout is planned. The routing for LHD haulage is shown in Figure 24.16. The design incorporates extraction drifts that are 5.0 m high x 4.6 m wide and spaced 30 m apart, and drawbell drifts that are similarly sized, aligned 60° to the extraction drifts, and spaced 18 m apart. The drawbells are created by drilling and blasting and once the drawbells connect to the undercut level, two drawpoints per drawbell are constructed. These drawpoints are supported with heavy support in the brows to withstand the attrition and erosion from the mineralized material as it is mucked. The resulting drawpoint spacing is 18 m by 15 m appropriate for the expected estimated fragmentation sizing of the mill feed.

 

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Figure 24.16Drawpoint Configuration for Kerr Underground

 

Source: WSP Golder (2022)

 

Overall, the mine plan assumes that to produce from Lifts 1 and 2 there will be 14 different levels and approximately 284 km of drifts and raises. Table 24.9 outlines the main drift types and dimensions used in the mine design.

 

Table 24.9Main Drift Types and Dimensions

 

Type Dimensions
Drawpoint and Extraction Drifts 5.0 m W x 4.6 m H
Perimeter, internal ramps 6.0 m W x 6.0 m H
Pre-conditioning, Undercut, Dewatering Drifts 5.5 m W x 5.5 m H
Access Tunnels, Conveyor Drifts 6.0 m W x 6.0 m H
Haulage Level 6.0 m W x 6.0 m H
Main Fresh/Return Air Drifts 7.0 m W x 7.0 m H

 

Vertical development will be constructed primarily using raisebore machines and raisebore-and-slash techniques. Table 24.10 summarizes the various types of vertical development and sizes.

 

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Table 24.10Vertical Development Dimensions

 

Type Dimensions Type
Intake Raises 3.5 m Diam. Bored Raise
Ore Passes 3.5 m Diam. Bored Raise
Exhaust Raises 2.4 m Diam. Bored Raise
Main Ventilation Raises 5 m Diam. Bored Raise
Ore Bins 5 m Diam. Bored Raise

 

Ventilation

 

The quantity of the required airflow was determined using a VentSim model and was based on the current mine design indicated earlier and 0.06 m3/s of ventilating air for each kilowatt of diesel-powered equipment. Mine equipment and engine utilization factors were also considered in developing the total ventilation needs.

 

The peak ventilation requirement was estimated to be 1,630 m3/s, which is sufficient to dilute noxious gases, particulate matter, and heat produced by the mining equipment and the operational activities on each mining level. Ventilation is provided by four main intake drifts and four main exhaust drifts 6 m wide and 7 m high. The second lift is ventilated via 3 fresh air raises and 4 exhaust raises of 5 m diameter which connect to the Lift 1 drifts, and to the fresh air main ramp and fresh air service ramp.

 

Heating of mine air in the winter months is included in the design and cost estimates, and this will be accomplished using mine heaters located at each of the two main fan installations. A mine air heating system of 18 MW capacity is required to heat 1,630 m3/s of air from an average minimum of -5.7°C to 2°C for six months per year (November through April). The average propane requirement is 6.5 million liters per year during production.

 

A ventilation schematic is shown in Figure 24.17.

 

Figure 24.17Ventilation Schematic

 

 

Source: WSP Golder (2022)

 

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Mine Dewatering

 

The water inflow rate for a 1 in 200-year storm event is predicted to be approximately 5 m3/s for Lift 1 and 9 m3/s at the completion of Lift 2. Other inflows that need to be dealt with are small quantities of groundwater and the production process water from drilling and dust suppression activities.

 

For Lift 1 most of the inflow water will be intercepted on the production level and routed away from the footprint to flow by gravity along the ventilation drifts to the portal at approximately 600 elevation. From here it will be directed into a 120 cm (48 inch) surface pipeline (separate from dewatering pipelines utilized for the Kerr open pit). This pipeline will convey the water to the WTP platform at 543 elevation by gravity and then it will be pumped up to the WSF. The 120 cm pipeline to the WTP is designed to handle

4 m3/s flow and any additional inflows that occur under rare extreme seasonal conditions will be treated as an emergency discharge.

 

For Lift 2, it is planned to intercept inflow water as it enters the production level and route it to sixteen 3,700 kW pumps (approximately 60,000 kW [80,000 HP]). These pumps have a capacity of 8 m3/s (65,000 gpm) to pump water up to the Lift 1 ventilation drifts from where it will flow under gravity to the ventilation drift portals. From the portals, as for Lift 1, water will be directed into two 120 cm surface pipes (an additional one added for Lift 2 mining). These pipes will convey the water to the WTP the same as for Lift 1. The estimated inflow rate over a 20-year period was used to determine an average annual pump utilization factor of 3.5% which was used to calculate the annual pump power requirements. The mine dewatering general arrangement is shown in Figure 24.18.

 

Figure 24.18Mine Dewatering General Arrangement

 

 

Source: WSP Golder (2022)

 

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2022 PEA Production Schedule

 

The proposed production schedule for Kerr, as developed in FF software, is shown in Figure 24.19. Note that the years shown apply to the overall KSM 2022 PEA mine production schedule. The pre-production construction will take 5 years with the first cave production being in PEA Year 7.

 

The ramp-up period is 5 years and is governed by the draw point opening rate of 120 per year and maximum draw rate of 80 m per year (220 mm per day). Both are consistent with industry experience. The total peak production rate of 80,000 tpd is achievable based on the available footprint area. Starting around PEA Year 25 there is a dip in total production as lift 1 depletes and Lift 2 commences. This lower production interval is the result of a 60-degree offset applied to the Lift 2 mining relative to Lift 1 to ensure that the Lift 1 crusher and associated access drifts on 625 m level are not impacted by caving from Lift 2. There is an additional 5 years (about 53 Mt) of ramp down production from the second lift that was removed from the mine plan so as not to run the mill at a reduced throughput, which would be an inefficient use of the mill asset.

 

Figure 24.19Kerr PEA Mine Production Plan

 

 

Source: WSP Golder (2022)

 

Only Indicated and Inferred Mineral Resources are included in the production plan. The production tonnes and grades are comprised of mineralized material included in the Mineral Resources reported in Section 14, mineralized dilution taken at grade (the portion of Indicated and Inferred material that is below the operating costs cut-off) and includes small percentage of material at zero grade. All the unclassified materials are treated as waste by setting the grades to zero. Table 24.11 shows the production tonnes and grade in the 2022 PEA mine plan based on their classifications.

 

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Table 24.11Kerr Production Tonnes and Grade in 2022 PEA Mine Plan

 

Lift 1 625RL Measured Indicated Inferred
Tonnes (millions) - 47.9 477.4
NSR (Cdn$/t) - 34.03 33.76
Au (g/t) - 0.25 0.30
Cu (%) - 0.53 0.48
Ag (g/t) - 1.3 1.7
Lift 2 130RL Measured Indicated Inferred
Tonnes (millions) - - 299.7
NSR (Cdn$/t) - - 35.76
Au (g/t) - - 0.33
Cu (%) - - 0.49
Ag (g/t) - - 1.8
Combined Measured Indicated Inferred
Tonnes (millions) - 47.9 777.2
NSR (Cdn$/t) - 34.03 34.53
Au (g/t) - 0.25 0.31
Cu (%) - 0.53 0.49
Ag (g/t) - 1.3 1.7
Note:Includes mineralized dilution taken at grade (the portion of Indicated and Inferred material that is below the operating costs cut-off) and includes small percentage of material at zero grade.

 

The mining sequence supporting the production is initiated on Lift 1 at a location where NSR values are the highest so that early value is maximized. The mining then propagates radially from this location and continues down the southern limb of the footprint. Lift 2 mining starts on the northern end and propagates southward to allow for concurrent production from both Lifts.

 

Development Schedule

 

The pre-production phase of the Kerr mine is estimated to be 5 years, which includes mine access, some initial footprint development, and construction of major mine infrastructure such as an underground crusher, material handling system, shops, dewatering system and primary ventilation. The development schedule was generated using Deswik software, and it assumes advance rates of 4 m/day and a maximum of 6 active headings in the first 2 years and 10 active headings thereafter.

 

The overall development strategy is to provide access to the planned crusher location as soon as possible and to prioritize the development of the service drift and MTT conveyor drift. Following this, extraction drifts, drawpoints, undercut, and pre-conditioning levels will be advanced to allow for cave initiation. Two intake and two exhaust airways will be established during pre-production with the remaining airways driven as production requires.

 

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Blasting and Explosives

 

Explosives will be received from suppliers and stored in the main surface explosive magazine. The bulk of the material will be ANFO for horizontal development and emulsion for up-holes, plus initiators and accessories.

 

Underground explosives magazines will be constructed for both Lift 1 and Lift 2 to service development and secondary reduction needs. Each magazine will have a one-week capacity.

 

Explosives will be used underground for five main purposes:

 

preparation for the development of mineralized areas

 

draw bell construction

 

undercutting

 

development in barren material

 

secondary blasting of boulders and hang ups

 

Mining Equipment

 

The peak primary and support equipment requirements for the Kerr mine are summarized in Table 24.12.

 

Table 24.12Peak Mobile Equipment Requirements for Kerr

 

Development

Equipment

Unit

Production

Equipment

Unit

Support

Equipment

Unit
Jumbo 12 LHD 21 Electricians Truck 3
Haul Truck 11 Truck 19 Grader 4
LHD 5 Drill Rig 5 Lube Truck 4
Bolter 15 Block Holer 7 Mobile Crane 2
Explosive Loader 4 Mobile Rockbreaker 4 Mine Ambulance 1
Scissor Lift 7 Water Cannon 4 Fire Truck 1
Shotcrete Sprayer 5 Boom Truck 1 Mine Rescue Truck 1
Transmixer 4     Backhoe 4
        Water Truck 1
        Sludge Truck 1
        Cable Bolter 4
        Telehandler 5
        Small Personnel Carrier 16
        Large Personnel Carrier 6

 

Mine Workforce

 

The estimated workforce for Kerr is approximately 529 persons at the peak of the pre-production construction period, and an average of 580 persons during the production period.

 

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24.16.5Mine Production Schedule (Open Pit and Underground Combined)

 

The summarized production schedule results are shown in Table 24.13 and Figure 24.20, including both open pit and underground mining. The mine production plan starts using conventional large-scale open pit mining equipment before transitioning into block cave underground bulk mining. The open pit at Upper Kerr, and the underground operations at Iron Cap and Lower Kerr are sequenced from higher grade and lower development and operating cost areas first (higher profitability to lower) to provide optimal cash flow. Upper Kerr pit with its lower preproduction and development capital cost is scheduled to provide first mill feed followed by the underground operations at Iron Cap and then Lower Kerr.

 

The production from the block caves will ramp up and then ramp down towards the end of their life. Lower Kerr block cave ramp up begins after Iron Cap ramp up is completed to avoid overlap of block cave development.

 

Table 24.13Summarized Production Schedule – Open Pit and Underground

 

 

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Figure 24.20KSM PEA Mill Feed Production Schedule

 

 

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24.17Recovery Methods

 

The proposed KSM plant will have a name plate capacity of 195,000 t/d. Peak mill feed production in the PEA plan is estimated to be 170,000 t/d. The process plant will receive mill feed from the Kerr and Iron Cap mines.

 

Based on the available information, metallurgical test work performed on various samples provides a reasonable indication of the mineralogical and metallurgical performance characteristics of the materials for this 2022 PEA. The process flowsheet developed for the KSM mineralization is considered appropriate for the current 2022 PEA given the nature of the mill feed and similar metallurgical performance of the samples from the KSM deposits.

 

Several metallurgical test programs have been carried out to assess the recoverability of copper, gold, silver and molybdenum values using the PEA flowsheet. Characterization and metallurgical test work on Iron Cap and Kerr samples are presented in the pertinent test work reports, which are summarized in Section 13.0. This test work examined grindability, concentration by flotation of copper and gold into a saleable concentrate, and the further recovery of gold and silver from gold-bearing sulphide materials (cleaner scavenger tailings and pyrite concentrate) by leaching.

 

Metallurgical performance projections from this test work are presented in Section 13.6.

 

Comminution test work results indicate that the samples from all the deposits are moderately hard for SAG and ball milling.

 

Flotation test work (batch and locked cycle) indicates that the mineralization is amenable for concentration into a saleable copper-gold concentrate with no significant penalty elements. Following flotation, cyanidation tests showed that it was possible to extract gold and silver by CIL leaching of gold-bearing tailing from the cleaner scavenger separation and pyrite concentrates derived from pyrite flotation of rougher flotation tailing. Preliminary trade-off studies indicate that cyanidation of Kerr and Iron Cap gold-bearing sulphide materials may not be economic due to elevated copper/gold ratios and low gold adsorption rates onto activated carbon. Additional test work to improve gold and silver recovery and economic trade-offs are required to optimize the process. Cyanidation of gold bearing pyrite materials are therefore not used in the 2022 PEA.

 

The metallurgical test results obtained from the various test programs have been used to predict plant metallurgical performance parameters for copper, gold, silver and molybdenum. The metallurgical performance projections of the five KSM mineralized zones are summarized in Section 13.6. In addition, work has been performed to determine the consumption of reagents in flotation and cyanidation and grinding media in comminution.

 

The process circuit incorporates three stage crushing, one stage milling, conventional flotation processes for the recovery of copper, gold, silver, and molybdenum.

 

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24.17.1Process Plant

 

The process plant is designed appropriately in accordance with the test work and consists of five separate facilities for the handling and processing of mill feed:

 

Underground primary crushing and a handling facility at the proposed underground blockcaving and crushing facilities, excluding the materials from upper Kerr pit which will be trucked and crushed on the surface at the Mitchell site

 

the crushed mill feeds will be transported by conveyor systems from the Mitchell pit area and the Kerr and Iron Cap underground blockcaving areas to train loading hoppers

 

train transportation of mill feed through the MTT to the process plant

 

the process plant includes:

 

oCoarse mill feed stockpiles

 

oSecondary crushing

 

oTertiary crushing

 

oBall mill grinding

 

oCopper-gold-silver-molybdenum bulk flotation, regrinding, 3-stage cleaner flotation and copper and molybdenum separation flotation if molybdenum grade is economical to recover

 

oPyrite flotation and regrinding

 

copper concentrate dewatering and handling

 

tailing delivery system to deposition in the TMF.

 

24.17.2flowsheet description

 

There will be three grinding and flotation lines, each line will be able to handle 65,000 t/d. When the mill feed rates are low, especially during the initial startup and late years, the process lines can be operated based on mill feed rate.

 

Comminution

 

The primary crushing facilities located at the Mitchell OPC will reduce the upper Kerr. Mill feed is trucked from the Sulphurets open pit mine site to the Mitchell OPC for crushing. Similarly, Mill feed from the Kerr open pit mine site is trucked to the Mitchell OPC for crushing to approximately 80% passing 150 mm using two 1,524 mm x 2,260 mm gyratory crushers. Kerr and Iron Cap mill feeds are separately crushed at the blockcaving and crushing facilities located underground and transported separately to the MTT train loading surge bins via different conveyor systems.

 

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Crushed mill feed is reclaimed from the train loading surge bins onto the MTT trains for transport to the main process plant site located at Treaty site, approximately 22 km northeast of the mine site.

 

Mill feed from the mine site through the MTT tunnel is delivered to two parallel mill feed stockpiles. In the early years, the upper Kerr material with supplement of the Iron Cap material will be main feeds to the mill. Starting from PEA Year 6, the mill feed will be supplied from Iron Cap and Kerr underground mines, the latter of which will introduce feed to the plant in PEA Year 7.

 

There are two coarse material stockpiles at the Treaty site. The stockpiles will be located at the exit portal of the MTT tunnel and will have a total live capacity of 180,000 t (90,000 t each stockpile). The coarse mill feed will be reclaimed and be further crushed by four cone crushers (three in operation and one on standby, as design) and then six HPGRs in closed circuit with vibrating screens.

 

The screen undersized material from the HPGR circuit will be fed to six ball mills in a closed circuit with hydrocyclones. The grinding circuit will employ six conventional ball mills in three process lines to grind the HPGR product to a particle size of 80% passing 125 µm to 150 µm. Each ball mill will be in closed-circuit with a cluster of hydrocyclones. The hydrocyclone underflow will gravity-flow to the ball mill feed chute, while the overflow will gravity flow to three copper-gold rougher flotation lines.

 

Flotation

 

The products from the primary grinding circuits will be fed into copper-gold/molybdenum rougher/scavenger flotation circuits, consisting of three operation circuits in parallel. The copper rougher flotation concentrates from the flotation circuits will be reground in two stages to a particle size of 80% passing, approximately 15 to 20 µm in the tower mills.

 

The reground rougher concentrate will then be upgraded in a cleaner flotation circuit with three stages of bulk copper cleaner flotation, producing a copper-gold or copper-gold/molybdenum concentrate with an average grade of approximately 25% Cu. Depending on the molybdenum content in the copper-gold/molybdenum bulk concentrate, the concentrate may be further treated by flotation to produce a molybdenum concentrate and a copper-gold concentrate. The molybdenum concentrate will be leached using the Brenda Mines procedure to reduce copper and lead contents.

 

The final copper concentrate will be dewatered by a combination of thickening and pressure filtration to approximately 9% moisture before being transported to the Stewart port site for ship loading and delivery to copper smelters, while the molybdenum bulk concentrate will be further dried prior to being shipped in bags to the port at Prince Rupert for delivery to molybdenum smelters.

 

The copper-gold/molybdenum rougher scavenger flotation tailing will be subjected to further flotation to remove remaining sulphide minerals, mainly pyrite, which will be disposed together with the cleaner scavenger tailing into a separate lined tailing storage cell in the TMF. Depending on the copper and gold contents, the pyrite rich flotation concentrate may be reground in tower mills to a particle size of 80% passing approximately 20 µm and floated for residual copper and gold values. The copper concentrate will be recycled back to the copper-gold bulk flotation circuit.

 

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The comminution circuits and flotation circuits are depicted in Figure 24.21.

 

Reagents

 

All the reagents will be prepared in a dedicated reagent preparation and storage facility within a containment area. Liquid reagents will be added in the undiluted form via metering pumps. Solid reagents will be prepared into adequate strength solutions in dedicated mixing tanks and stored in holding tanks to be added to the processes via metering pumps.

 

Water and Pressurized Air Supply

 

Three separate water supply systems will be provided to support the operation:

 

A freshwater system.

 

A process water system for grinding/flotation circuits.

 

Plant air service systems will supply blower air to flotation, and high-pressure air to filtration and general plant and instrumentation services.

 

Assay and Metallurgical Laboratory

 

The assay laboratory will be equipped with necessary analytical instruments to provide routine assays for the mine, process, and environmental departments.

 

The metallurgical laboratory, with laboratory equipment and instruments, will undertake all necessary test work to monitor metallurgical performance and to improve the plant production and metallurgical results.

 

Process Control and Instrumentation

 

The plant control system will consist of a DCS with PC-based OIS located in control rooms at the process facilities. Process control will be enhanced with the installation of an automatic sampling system. The system will collect samples from various lines for on-line analysis and the daily metallurgical balance.

 

CCTV support will be provided at various locations at the crushing and plant facilities to ensure comprehensive site monitoring.

 

Process Flow Diagram

 

Figure 24.21 below illustrates the overall process block flow diagram. The major design consideration in the process plant equipment sizing and layout is the use of the largest equipment sizing available in order to minimize pumping and piping requirements, process building footprint and capital costs.

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Figure 24.21Overall Process Block Flow Diagram

 

Source: Tetra Tech (2022)

 

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24.18Project Infrastructure

 

24.18.1On-Site Infrastructure

 

The water and waste management structure designs in the 2022 PEA are based on those described in the approved EA with conceptual-level adjustments to the 2022 PEA mine plan layouts, staging plan, and capacities required for the 2022 PEA mine production schedule.

 

The approximate length of existing single axis water diversion tunnels at PEA Year 1 in the 2022 PEA is 18 km and comprise the MDT, MTDT, NPWDA and MVDT.

 

Rock Storage Facility

 

The 2022 PEA rock storage requirements are relatively small considering there is only one deposit mined by open pit methods producing waste rock (e.g., Upper Kerr). The 2022 PEA waste management plan consolidates all waste rock into a single facility, including minor waste volume from initial underground development, into the Mitchell Valley as pit backfill which is contiguous with the Mitchell RSF. No waste rock in the 2022 PEA is placed in the McTagg RSF or the Sulphurets Pit Backfill RSF.

 

Water Management

 

The overall 2022 PEA operational water management and water treatment designs are, aligned with management plans described in the approved EA and no adjustments to the EA boundary were required since the three areas to be mined in the 2022 PEA mine plan were inside disturbance areas identified in the EA.

 

The WSD crest elevation is kept at the same elevation as in the approved EA for the 2022 PEA. With the reduced tonnage of waste rock, WSD storage and WSD crest elevation could potentially be reduced, although these optimizations have not been adopted for the 2022 PEA.

 

HDS Water Treatment Plant

 

In the 2022 PEA the catchments were reduced from the EA due to a reassessment of catchment areas, a change in pit areas, and reducing the total waste rock tonnage. This reduction in catchment area has resulted in a reduction in the amount of run off from the catchment area that has to be treated prior to discharge to the environment (KCB, 2020a). The direct effect of this reduction is that the design rate for the HDS water treatment plant (freshet maximum) that processes the mine water run-off, resulting in a reduction in the estimated number of HDS water treatment trains required from seven for the EA to about four for the 2022 PEA at the concept level.

 

Tailings Management

 

TMF management philosophy and criteria for the 2022 PEA are generally consistent with the TMF designs and configuration presented in the approved EA. The 2022 PEA has a peak throughput with an annual average of 170,000 t/d during early operations, so tailings storage would account for this placement rate and maintain requisite supernatant pond operational volume and freeboard where storage occurs.

 

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Cost allowances for tailings storage and management have been included in the 2022 PEA. Tailing management is envisioned as a combination of technically viable storage approaches that can be refined in future studies that can comprise appropriate and responsible solutions depending on best selected technology. Early definition of extended tailings management beyond the permitted capacity would be required well before deciding to proceed with the 2022 PEA production plan to allow for exploring all facets of best available technology.

 

MTT Rail System

 

MTT rail system will transport crushed mill feed, freight, fuel and personnel between the Mitchell OPC and Treaty OPC.

 

Crushed mill feed will be conveyed to one of three 15,000 t capacity underground bins. Loading chutes under the bin will feed mineralized material into train wagons for transport to Treaty OPC where the wagons will bottom dump the crushed mill feed into a 15,000 t capacity surge bin. Apron feeders will then reclaim the mill feed to a belt conveyor which will report to the 60,000 t live capacity COS.

 

Ten mill feed trains will be needed to deliver an average of 170,000 t/d of mill feed to the process plant. Each mill feed train will be made-up of two 140 t electric locomotives and 16 of 42 m3 bottom-dump mill feed wagons.

 

Specially configured personnel and freight trains will transport freight, fuel, and personnel through the MTT. Staging areas at each end of the MTT for marshalling and loading/unloading trains at each end will separate these activities from those for mill feed transport. Personnel, freight, and fuel handling will be scheduled during the day shift operations. A maintenance shop and siding for rolling stock will be located in the Treaty staging area.

 

Train operations, with the exception of train loading, will be controlled by an automated controls and scheduling system. The main control room will be located in the Treaty staging area. Locomotives will be unmanned and not require engine drivers.

 

The MTT rail system can increase short term upside capacity when required for production recovery by adjusting track time allocation between mill feed, freight and personnel transport. Longer term recovery requirements will need to be addressed in future studies with considering Treaty stockpiling, detailed track logistics studies, and rolling stock availability.

 

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Facilities, Buildings, and Services

 

Facilities, buildings and services in the mine site and Treaty OPC will include the following:

 

site access roads

 

fuel receiving, storage, and dispensing

 

sewage treatment

 

medical buildings and ambulance stations

 

administration offices

 

consulting engineers’ and contractors’ offices and storage/laydown areas

 

warehouses, cold storage, and maintenance buildings

 

truck maintenance and wash bay

 

batch plants

 

accommodation facilities including receptions, lunch rooms and cafeterias, mine dry and wash cars, recreation facilities, kitchen, and parking areas

 

potable water treatment

 

potable, process, fire and fresh water systems

 

container storage

 

laydown and equipment and materials storage areas

 

landfill

 

waste management storage and handling and Incinerators

 

security fencing and gates

 

helipads

 

explosives magazines and AN prill storage areas

 

explosives manufacturing plant

 

communication systems

 

auxiliary truck and support vehicles

 

electrical power supply and distribution.

 

24.18.2Off-site Infrastructure

 

24.18.2.1Electrical Power Supply

 

Electrical power will be supplied to the KSM Site from the BC Hydro NTL 287 kV transmission line that runs from the Skeena Substation near Terrace, BC, to a substation near Bob Quinn Lake, further north of the KSM Site.

 

KSM will connect to the NTL transmission line at the new Treaty Creek Terminal (Switching Station) located adjacent to Highway 37, approximately 18 km south of Bell 2 Lodge. This station, designated TCT by BC Hydro, is currently under construction. A new 30 km long, 287 kV transmission line tap will connect this substation to the KSM Main Substation No. FLT1 at the Treaty OPC.

 

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BC Hydro has completed an updated System Impact Study (SIS) that confirms a Contract demand of 245 MW for KSM. KSM has signed (Feb. 2022) a Facilities Agreement for the supply of 25 MW of construction utility power supply. A second Facilities Study will be carried out during the KSM initial construction period leading to a second Facilities Agreement for the supply of 245 MW of power when mine operations start.

 

The cost of electric power at the site main substations, including transformer and line losses, is Cdn$0.0596 per kilowatt hour.

 

The KSM 287 kV transmission line tap will terminate at the compact gas insulated (GIS) main substation designated FLT1 at the Treaty OPC. This substation steps the voltage down to 25 kV to supply the various process plant and other area feeders and 138 kV to supply the cable feed thought the MTT to the Mitchell area.

 

The 25 km long 138 kV cables through the MTT will terminate at GIS substation FLT2 located at the Mitchell OPC. This station steps the voltage down to 25 kV for local distribution and 69 kV for distribution to more remote facilities. The substation has an installed capacity of 270 MVA with two stages of transformer fan cooling in operation (under emergency conditions) and provides full redundancy of electric power supply and flexibility for maintenance.

 

The 2022 PEA includes 3 energy recovery systems. These systems recover energy from the McTagg diversion tunnel, the discharge from the WSF to the WTP, and the North Tailings Line. The plants generate an average of 60,122 megawatt hours per annum over the first 10 years of operation with an annual value of Cdn$6,078,000, based on the BC Hydro Tier 2 rate for energy. Subsequent to the first decade, the generation varies somewhat, but the McTagg and WTP energy recovery systems will continue to operate after mine closure.

 

24.18.2.2Access Roads

 

The access roads to the KSM site during the 2022 PEA will be as follows:

 

Eskay Creek Access Road

 

CCAR

 

TCAR

 

Upper NTAR

 

24.18.2.3Logistics

 

Inbound equipment and materials will be transported either by barge to Stewart, BC, or by rail to Terrace, BC, where these loads will be consolidated at local marshalling yards or staging areas for onward transport by truck to KSM.

 

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Copper concentrate will be transported by truck to a deep water port in Stewart, BC, where it will be held in storage until loaded onto oceangoing vessels.

 

24.18.3Iron Cap Project Infrastructure

 

Road and Logistics

 

The main access for personnel and materials to the Iron Cap operations is through the MTT tunnels originating in the Treaty Valley and connecting to the Mitchell OPC. Access is then via the Iron Cap access tunnel collared nearby. All personnel will stay in a mining camp located in the Mitchell Valley area. Materials required for the operation of the mine will be delivered via the MTT tunnels to warehouses in the Mitchell Valley.

 

Underground Crushers and Conveyors

 

The Iron Cap block cave will produce 90kt/d at steady state production. The mill feed will be transported by LHD directly from the drawpoints to one of three ore passes located along each extraction drift. Each pass is fitted with a stationary rock breaker and the sized material gravity feeds to the haulage level where continuous loading chutes transfer the uncrushed material into train cars. Electric locomotives then transport the cars to one of two crusher unloading stations where the material is bottom dumped into the crushers. Two 60 in x 75 in gyratory crushers will be installed to crush the mill feed to 152 mm size.

 

The crushed material will be transported by a 2.4 km long series of conveyors feeding two ore bins that in turn feed the MTT trains for transfer to the process plant. The conveyors will vary in width from 54 in to 72 in depending on location and requirement. The belts do not exceed a 15% grade.

 

Conveyor structures will be back mounted to provide ease of cleaning under the belts. Conveyor drives and tail and head pulleys will be sill mounted for ease of installation and maintenance.

 

Underground Mine Service Water System

 

The mine will require 199 m3/h of process water for the bolters and for the development and production drills. In addition, it is estimated that 50 m3/h of water will be required for the conveyor fire suppression systems. This water will be supplied through the Iron Cap access tunnel from pumps in the Mitchell Valley. The water will be delivered to the working face through steel pipes.

 

Underground Electrical Reticulation System

 

The Iron Cap mine will require approximately 24.4 MW of average electrical power at peak operation. The main contributors to this demand are the crushers, conveyors, ventilation fans, and drill jumbos. The annual power consumption is estimated to be 213,700 MW hours per annum during peak production.

 

The main power will be supplied to the underground from Mitchell area substation FLT2 through two redundant 25 kV cables hung from the back of the Iron Cap access tunnel and MTT train tunnel to create a ring main style system. Each of the main levels will have a 25 kV feed except the haulage level will have separate feeders for the west and east train loops and the extraction level will also have two separate feeders. The 25 kV distribution voltage will be stepped down to the required utilization voltage by skid mounted dry-type unit substation transformers. Equipment that draws larger loads (e.g., ventilation fans, conveyors and crushers) will be fed from local permanent transformers.

 

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Waste Storage Facilities

 

Approximately 4.8 Mt of development rock (mineralized and non-mineralized) will be generated during the four-year pre-production period and hauled by truck to the Mitchell OPC. The 2022 PEA mine plan assumes the NPAG portion of this material will be used in construction of pads and the other material will either be hauled to the Mitchell RSF or fed into the Mitchell mill feed. After PEA Year 1, once one of the crushers and the conveyor are installed, all development material will be crushed and conveyed to the MTT bins and hauled to the mill by the trains. Most of this material will be mineralized rock from within the footprint.

 

Water Management

 

The underground mine dewatering system (refer to 24.16.3 Mine Dewatering) is discharged into the North Pit Wall Dewatering Adit (NPWDA) via graded drifts that allow for water to drain by gravity. From there, it will flow by gravity to the WSF.

 

Surface Infrastructure

 

The 2022 PEA design assumes the surface infrastructure required to support the Iron Cap operations will be integrated with the overall site infrastructure. This will include major equipment maintenance and fuel provision.

 

Surface infrastructure required to support operations include:

 

equipment shop for major repairs

 

explosives magazine

 

warehouse

 

concrete plant

 

water pump stations

 

general mine offices

 

control room and training center

 

first aid station and mine rescue room

 

construction and operations camp

 

mining contractor facilities

 

change house

 

instrumentation shop

 

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Underground Infrastructure

 

Underground facilities at Iron Cap include:

 

equipment shop

 

warehouse

 

fuel bays

 

explosives magazines

 

sampling room

 

refuge stations (fixed and portable)

 

lunchrooms

 

offices, map and meeting room, training room

 

first-aid station

 

comfort stations

 

water tanks

 

fire water tanks and fire suppression systems, primarily for crushers, conveyor transfers and shops.

 

24.18.4Kerr Project Infrastructure

 

Road and Logistics

 

The main access for personnel and materials to the mining complex in Mitchell Valley is through the MTT tunnels connecting the Treaty Valley to the Mitchell OPC. From the Mitchell OPC, a series of roads developed for the overall KSM site will lead to the Kerr surface complex area just outside of the Kerr portals. All personnel will stay in the mining camp located in the Mitchell Valley area. Materials required by the Kerr mine will be delivered via the MTT tunnels to the warehouses in the Mitchell Valley.

 

Kerr mine is accessed through a series of portals and ventilation adits along the northwestern base of Kerr Mountain. Mine access is primarily through two of the intake drifts, one of which directly connects to the footprint area as a main ventilation intake drive, and the other eventually parallels the UG conveyor as a service drive that leads to the crusher area. These two primary drifts will connect via ramps to the lower lift (Lift 2) to provide access to the lower elevation.

 

Underground Crushers and Conveyors

 

The Kerr block cave will produce 80 kt/d at steady state production. The mined material will be transferred from production drawpoints to mineralized material passes fitted with stationary rock breakers to size the material before it gravity feeds to the haulage level below. At the haulage level, 55 t trucks will haul the material to the crushers.

 

The mine will produce from two cave Lifts over LOM, the 625 level (Lift 1) and the 130 level (Lift 2). Each lift will be serviced by two, 60 in x 75 in gyratory crushers on the haulage level. These crushers are sufficient to support the peak production rate of

80 kt/d. Each crusher will have three dumping points for rear dump articulated underground haul trucks. The crushers will discharge to incline conveyor systems which connect to a series of conveyors to transfer the material to the MTT transfer bin at the portal.

 

There will be a total of 10.2 km of underground conveyor belts with widths ranging from 54 in to 72 in. Conveyor structures will be back mounted to provide ease of cleaning under the belts. Conveyor drives and tail and head pulleys will be sill mounted for ease of installation and maintenance. Most conveyor inclinations are at 15% or less and none exceed 17%.

 

Underground Mine Service Water System

 

It is estimated that the mine will require 203 m3/hr of process water for the bolters and the development and production drills. In addition, it is estimated that 50 m3/hr will be required by the conveyor fire suppression systems. Mine service water is sourced from the surface complex near the Kerr portal area and piped underground through a series of pipes to the main areas for each lift. Smaller pipelines will branch off the main feeder pipe on each level to support excavation of mine development headings, longhole drilling, shop activities, etc.

 

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Underground Electrical Reticulation System

 

The Lower Kerr mine will require an average of 36.6 MW electrical power at peak operation. The main contributors to this demand are the crushers, conveyors, ventilation fans, and drill jumbos. This load is larger than typical as it includes requirements for significant pumping plus the long conveyor to the MTT portal. The annual power consumption is estimated to be 318,000 megawatt hours per annum during peak production.

 

The main power will be supplied to the underground from the Mitchell area substation

FLT2 via a 69 kV distribution line to a modular 69 to 25 kV surface substation near the mine. A 25 kV ring main system will feed the mine via Teck cables from this substation. Each of the main levels will have a 25 kV line which will be stepped down to the required voltage by skid mounted dry-type transformers. Equipment that draws larger loads (e.g., ventilation fans, conveyors and crushers) will be equipped with a permanent transformer.

 

Waste Storage Facilities

 

Approximately 7.3 Mt of waste rock (mineralized and non-mineralized) will be generated and brought to surface during the initial 5-year pre-production period. NPAG portion of this material will be hauled by truck to the Mitchell RSF. The mineralized material will be hauled to the Mitchell OPC and fed into the mill feed.

 

After 2022 PEA Year 7, once one of the crushers and the conveyor to the MTT are installed, all development material will be crushed and conveyed to the MTT bins and hauled to the mill by the trains. Most of this material will be mineralized rock from within the footprint.

 

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Water Management

 

The underground mine dewatering system (refer to Section 24.16.4 Mine Dewatering) is discharged to the surface complex established outside of the Kerr portal areas. From there the water will gravity flow via pipeline to the WTP platform and then be pumped to the WSF in the Mitchell Valley.

 

Underground Electrical Reticulation System for Dewatering

 

The mine dewatering system at Kerr requires an average of 4 MWh with a maximum of 30 MWh during a peak storm event, which is greater than the power requirements of the mine under normal conditions. The strategy during a peak storm event will be to shut down or reduce operations in the underground mine along with other site facilities when the high-powered pumps are required. This will allow power to be diverted from normal operations to power the pumps.

 

The main power will be supplied to the underground from the Mitchell area substation FLT2 through a 25 kV cable hung from the back of the access ramp and conveyor tunnel to create a ring main style system. Each of the main levels will have a 25 kV line which will be stepped down to the required voltage by skid mounted dry-type transformers. Equipment that draws larger loads (e.g., ventilation fans, conveyors and crushers) will be equipped with a permanent transformer.

 

Surface Infrastructure

 

The 2022 PEA design assumes the surface infrastructure required to support the Lower Kerr operations will be integrated with the overall site infrastructure.

 

Surface infrastructure required to support operations will include:

 

explosives magazine

 

warehouse

 

water pump stations

 

concrete plant

 

general mine offices

 

control room and training center

 

first-aid station and mine rescue room

 

construction and operations camp

 

mining contractor facilities

 

change house

 

instrumentation shop.

 

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Underground Infrastructure

 

Underground facilities at Lower Kerr include the following on each lift:

 

extraction level equipment shop

 

haulage level equipment shop

 

warehouse

 

fuel bay

 

explosives magazine

 

concrete / shotcrete slickline and loading system (for Lift 2 only)

 

sampling room

 

refuges

 

lunchrooms

 

offices, map and meeting room, training room

 

first-aid station

 

comfort stations

 

service water tanks

 

fire water tanks and fire suppression systems, primarily for crushers, conveyor transfers and shops.

 

24.19Market Studies and Contracts

 

COPPER CONCENTRATE

 

Seabridge engaged NSA to provide opinion reports on marketing inputs in 2016 and review the 2019 copper-gold concentrate market and related concentrate treatment charges, excluding offsite transportation costs. In May 2022, NSA was asked to provide an update to this opinion focusing on the validity of the assumptions. The current information and options mainly come from the 2016 and 2019 Opinion Reports with applicable comments where appropriate. The information and options in this section come from the opinion reports. No smelter contracts are currently in place or being negotiated. All currency amounts used in this section are in US dollars, unless otherwise specified.

 

MARKETABILITY

 

When considering the marketability of copper concentrates, quality and quantity are determining factors. There are considerable variations in the quality of concentrates and the requirements of various smelters do vary; such variations relate to the technical abilities of the smelter and its overall concentrate feed and blend. Gold, silver and copper are key elements in determining concentrate salability, as well as iron, sulphur and other impurities.

 

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While smelters prefer a feed with about 30% copper and similar amounts of iron and sulphur, copper grades from many major high-grade suppliers have been falling and the market is seen the blend for many smelters dropping to a copper content of about 27%. Apart from the copper content, levels of iron, and sulphur, other key elements including gold and silver and impurity content are factors in concentrate market.

 

Based on the impurity levels projected by Tetra Tech using the test results completed to date, concentrates are relatively clean. Depending on the prevailing market at the time of contract negotiations, penalties will likely be minimal if any. Certain smelters in Japan, South Korea, and Europe, have more interest in copper concentrates with high gold content.

 

COPPER CONCENTRATES SMELTING TERMS

 

NSA suggests that the annual benchmark terms are likely to be a guide to future levels. for 2022, a copper treatment charge with a range being TC $90 to $95/dmt and RC $0.090 to 0.095/lb of refined copper in constant dollars can be assumed. The copper-gold concentrate market review by NSA in Q4 2019 foresaw that the treatment charges and refining charges would be slightly lower than the 2016 projections. For 2022 PEA, Tetra Tech suggests keeping the smelting terms unchanged. Also, the payable terms identified in 2016 should be kept unchanged as below.

 

PAYABLE METALS

 

  Copper Pay 96.5% with a minimum deduction of 1 unit (amount deducted must equate to a minimum of 1% of the agreed concentrate copper assay).
   
  Silver If over 30 g/dmt pay 90%.
     
  Gold A scale is applicable with some variations of the following:

 

less than 1 g/dmt, no payment

 

1 to 3 g/dmt, pay 90%

 

3 to 5 g/dmt, pay 93%

 

5 to 7 g/dmt, pay 95%

 

7 to 10 g/dmt, pay 96.5%

 

10 to 20 g/dmt, pay 97.0%

 

over 20 g/dmt pay 97.5%

 

over 30 g/dmt pay 97.75%.

 

REFINING CHARGES

 

  Copper: $0.095/lb payable copper
   
  Gold: $6.00 to $8.00/oz payable gold

 

Silver:$0.50/oz payable silver

 

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TREATMENT CHARGES

 

Treatment Charge:$95.00/dmt CIF-FO main smelter port

 

PRICE PARTICIPATION

 

Not applicable at present.

 

PENALTIES

 

Arsenic:$2.50 to $3.00 per 0.1% over 0.1% up to 0.5% arsenic

 

Antimony:$3.00 to $4.00 per 0.1% over 0.1% antimony

 

Lead:$2.00 to $3.00 per 1% over 0.5% to 1.0% lead

 

Zinc:$2.00 to $3.00 per 1% over 2% to 3% zinc

 

Mercury:$2.00 per each 10 ppm over 10 ppm mercury

 

Bismuth:$3.00 to $5.00 per 0.01% over 0.03 to 0.05% bismuth

 

Selenium:$3.00 to $5.00 per 0.01% over 0.05% selenium

 

Tellurium:$4.00 to $5.00 per 0.01% over 0.02% to 0.03% tellurium

 

Fluorine:$1.00 to $2.00 per 100 ppm over 300 ppm fluorine

 

Chlorine:$1.00 to $3.00 per 100 ppm over 300 ppm chlorine

 

Furthermore, penalties may also vary from smelter to smelter. It should be noted that for the elements where a percentage range is used, this relates to ranges of penalty thresholds that are negotiated.

 

OTHER OFFSITE COSTS

 

Various indirect costs other than smelter charges include:

 

losses are assumed to be 0.1% or less

 

marine transport insurance is assumed to be in the range of 0.10 to 0.15% of net invoice value of the concentrate

 

third party supervision, assaying and umpire costs are assumed to be approximately US$1.00/dmt

 

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the cost of marketing varies with concentrate tonnage, location, and number of smelters to be shipped. For the 2022 PEA, the estimated marketing cost is in the range of US$5.00 to US$10.00/dmt

 

the copper concentrate transportation costs are based on the following assumptions by Tetra Tech:

 

otrucking: US$35/wmt

 

oport storage and handling: US$17/wmt

 

oocean transport to Asian port: US$49.00/wmt.

 

24.20Environmental Studies, Permitting and Social or Community Impact

 

The KSM mine development plan was subject to the BC Environmental Assessment Act (BCEAA, the Act), the Canadian Environmental Assessment Act- 1992 (CEAA), and Chapter 10 of the Nisga’a Final Agreement (NFA).

 

As of May 2022, the Property has completed the provincial and federal environmental assessment review processes, in accordance with the principles of the Canada-BC Agreement on Environmental Assessment Cooperation (Cooperation Agreement 2004), and the appropriate certificates/approvals have been obtained for this stage of the KSM Property’s development. Additionally, permits for early-stage construction activities, continuation of exploration, certain permit and project approval renewals have also been obtained.

 

KSMCo has developed long term respectful relationships with the Nisga’a Nation, the Tahltan Nation and the three other First Nations groups who are potentially influenced by the KSM development, over the past twelve years. These relationships, including the requirements contained within the Benefits Agreement signed with the Nisga’a Nation in June 16, 2014, the Impact Benefit Agreement signed with the Tahltan in July 8, 2019 and the Sustainability Agreement negotiated with the Gitanyow Wilps, also signed in June 16, 2014 respectively, would remain in good standing and the agreements would continue to be adhered to by the Property operating company if the proposed mine plan outlined in the 2022 PEA is implemented.

 

KSMCo has undertaken comprehensive biophysical and socioeconomic studies and reports in support of the Application /EIS (Rescan 2013) in relation to the EA process and permitting, and continues specific monitoring and studies to comply with approval conditions, management plans and further permitting. Fish habitat offset plans are being permitted and constructed. The 2022 PEA covers the same geographic area and valued environmental and social components, and will involve similar disturbances required to develop infrastructure and the mine as was previously assessed and approved.

 

Based upon completion and evaluation of additional technical studies required to identify and examine the net environmental benefit of the 2022 PEA, regulatory approval would be forthcoming only after the appropriate engagement and information sharing has occurred with the Nisga’a Nation, the Tahltan Nation and First Nations whom have an interest in KSM.

 

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The 2022 PEA mine design works within the EA to reduce the overall impact of developing the Kerr and Iron Cap deposits and the following items:

 

large majority of mill feed produced from underground block cave mines with a subsequent smaller surface disturbance relative to open pit mining

 

placement of the proposed tonnage of waste rock from Upper Kerr open pit mining on prior disturbed areas in Mitchell Valley from mining activity depicted in the 2022 PFS

 

utilizing infrastructure built during the 2022 PFS production plan, thus no major additional disturbance over that plan and maintaining impact within the limits of the EA.

 

An extensive system of water management facilities will be constructed and maintained throughout the life of the KSM mine to divert fresh (non-contact) water away from disturbed areas and to collect water that has contacted disturbed areas (contact water) for treatment before release into the environment

 

The proposed management of waste rock is outlined in the Rock Storage Facilities Management and Monitoring Plan, which can be found in Volume 26 of the Application/EIS (Rescan 2013), in accordance with definition as a major dump as defined under Section 10.5.5 of the Code (BC EMLI 2021).

 

Mine reclamation and liability cost estimates are developed in accordance with the Interim Major Mine Reclamation Security Policy (EMLI 2022), Code and applicable guidelines, The closure plan costs reflect the mine plan outlined in the 2022 PEA. Closure costs comprise physical closure (e.g., grading, covering and revegetation); access costs for post-closure activities; monitoring, inspection, and maintenance; and post-closure water treatment.

 

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24.21Capital and Operating Cost Estimates

 

The capital cost estimate is the product of engineering supporting the 2022 PEA. The unit rates used for the 2022 PEA capital and operating estimates (labour, power, fuel and other consumables) are based on recent price quotes and estimates. Well established internal benchmarks were used to check capital and operating costs to verify reasonableness of cost levels. These benchmarks are internal to each company.

 

The capital cost estimates were produced by the consulting firms named in Table 24.14 with the area of responsibility for each firm identified.

 

Table 24.14Capital Cost Estimate Responsibilities by Firm

 

Capital Cost Area Firm Responsible
Open pit mine development and related infrastructure, equipment, dewatering, wall depressurization, capitalized operating costs, MTT and associated infrastructure, indirects, and contingency MMTS
Underground mine development and related equipment, dewatering, ventilation and services, infrastructure, indirects, and contingency WSP Golder
Underground material handling for Iron Cap and Kerr block caves Tetra Tech
Process plant and related ancillary facilities Tetra Tech
Electrical power supply and energy recovery W. N. Brazier Assoc.
TMF KCB
Construction indirects (except for mining, TMF), and contingency (except for mining, TMF) Tetra Tech

 

24.21.1Capital Cost Estimate

 

The 2022 PEA level estimate includes:

 

direct field costs of executing the mining, construction, installation and commissioning of all structures, utilities, materials, and equipment

 

indirect costs associated with spares, initial fills, freight/logistics, commissioning, EPCM, and vendor assistance

 

contingency.

 

See Table 24.15 for a summary of the 2022 PEA capital cost estimate, summarized by activity.

 

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Table 24.152022 PEA Capital Cost Estimate Summary

 

Area Initial Sustaining Total
US$ M US$ M US$ M
Direct Costs      
Mine 828 6,678 7,506
Process 0 651 651
TMF 74 664 738
On-site Infrastructure 26 573 599
Power Supply/Energy Recovery 0 112 112
Total Direct Capital 927 8,678 9,606
Indirect cost 253 1,249 1,502
Contingency 320 2,824 3,145
Total Estimated Total 1,500 12,752 14,252

 

Notes:1 Sums may not add due to rounding
 

2 PST is not included, please see section 24.22 for details.

3 All costs associated with closure cost are addressed separately in financial analyses outside of capital.

4 Most of the sustaining on-site infrastructure is included in the mine category.

5 Mine direct costs include for crushing, conveying and power infrastructure.

 

Basis of Estimate

 

This estimate falls under the AACE® Class 5 Estimate classification and its accuracy is expected to be within -30% to +50% of final 2022 PEA cost.

 

Initial capital cost is defined as all costs associated with development of the operation until first mill feed in PEA Year 1. It includes mine, TMF and on-site infrastructure.

 

Sources of information for the 2022 PEA capital cost estimate include, but are not limited to, vendor quotations, in house database or historic data.

 

Indirect Costs and Contingency

 

Allowances were made for indirect costs based on in-house database, historic data or 2022 PEA parameters. The overall factor applied to the direct capital costs to estimate indirect costs was approximately 26.1% for non-mining categories.

 

An overall contingency of about 28.3% was applied for the 2022 PEA. Contingency is applied to the sum of direct and indirect costs.

 

Sustaining Capital

 

The majority of the sustaining capital costs comprise Kerr open pit, and development of the Iron Cap and Lower Kerr block cave mines.

 

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The sustaining capital for the Mine Site is US$6,678 billion. This number covers the direct capital costs for the LOM and includes all open pit and underground mining operations, as well as Mitchell pit backfill direct capital cost.

 

At the plant site, the majority of the sustaining capital cost is for replacing infrastructure at regular intervals during the PEA 39-year mine life.

 

Funding of Initial Capital

 

The 2022 PEA is evaluated as a standalone development after completion of the 2022 PFS mine plan. The owner would likely plan the development of the 2022 PEA over the final 4 years of the mine operations described in the 2022 PFS, however, the costs and revenues for the two studies are reported separately. The post-tax cashflow estimated for the last 4 years of the 2022 PFS significantly exceeds the estimated 4 years of initial capital for the 2022 PEA and the owner could fund construction of the 2022 PEA development from operating cashflow.

 

24.21.2Operating Cost Estimate

 

The operating cost estimate is the product of engineering developed during the previous studies and the 2022 PEA study. The unit rates used for the 2022 PEA estimates (labour, power, fuel and other consumables) are based on recent price quotes in 2022 Q1/Q2 and estimates from the consultant’s databases.

 

The 2022 PEA operating cost estimates were produced by consulting firms named in Table 24.16 with the area of responsibility for each firm identified.

 

Table 24.16Operating Cost Estimate Responsibilities by Firm

 

Operating Cost Area Firm Responsible
Open Pit Mining MMTS
Underground mining WSP Golder
Processing, G&A and site services TT
Tailings disposal KCB
HDS water treatment Catchment area by KCB and costing by TT
Se Water Treatment Quantities by BQE Water Inc. and costing by TT
Energy recovery W. N. Brazier Assoc.
PST TT

 

Basis of Estimate

 

Operating costs detailed in the 2022 PEA were derived from a variety of sources including, but not limited to, benchmarking analysis and factoring costs from previous studies where possible.

 

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The operating cost estimate is considered to have a level of accuracy of –25% to +35%. Overall assumptions for operating costs:

 

costs are presented in 2022 US dollars, unless stated otherwise. When required, certain costs in this report have been converted using a fixed currency exchange rate of Cdn$1.00 to US$0.77

 

the costs per dry metric tonne of material milled (US$/dmt) provided in this report are the average costs over the LOM.

 

The average LOM operating cost per tonne milled for the 2022 PEA is estimated at US$11.98/dmt. Details are presented in Table 24.17.

 

The LOM average unit operating cost is calculated by dividing the total LOM operating costs by the LOM dmt milled. The costs exclude pre-production mining, reclamation and fund costs for servicing long-term water treatment that are accounted for separate from operating costs. The operating cost estimates also include:

 

LOM annual energy recovery revenue credit of US$-0.09/dmt derived from on-site power generation from hydro energy recovery systems, and

 

life of mine average provincial sales tax (PST) which is estimated at US$ 0.05/dmt.

 

Table 24.17 2022 PEA Average Operating Costs

 

Operating Costs Cost Cost
(US$/dmt milled) (LOM US$ million)
Mining (OP & UG)  4.99 8,451
Milling /Train Transport  4.31 7,292
Tailings 0.15 249
G&A  1.39 2,345
Site Services  0.50 847
Water Management 0.68 1,155
Annual Energy Recovery (0.09) (151)
PST 0.05 85
Operating Costs/dmt milled 11.98 20,273

 

Open Pit Mine Operating Costs

 

Open pit operating cost were estimated using an Excel® based cost model utilizing first principle build up for the haulage costs and benchmarking from a similar study for all other cost areas. The most significant components of the operating costs are fuel and labour. The average LOM cost for the open pit mine is US$2.85 per tonne mined. On a per tonne milled basis, the cost is US$7.19/dmt. (see Table 24.18). Open pit mining represents approximately 7% of the mill feed source.

 

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Table 24.18Open Pit Mine Operating Costs

 

Open Pit Mining Operating Costs Units Cost
Total Cost  US$ million 891
Cost per ex-pit tonne mined  US$/t 2.85
Cost per mill feed tonne milled  US$/dmt 7.19

 

Underground Mine Operating Costs

 

The major components of the underground mining cost are labour and mobile equipment maintenance. Operating costs include the following indirect costs:

 

mine management, supervision, technical staff

 

freight on bulk materials

 

PPE and training.

 

Underground mining costs were estimated for each block cave operation using similar methodology as used for the capital cost estimate. The LOM cost per tonne mined varies between US$4.34 (Iron Cap) and US$5.25 (Kerr). The Iron Cap mining costs are lower due to the application of battery electric loaders, electric haulage trains and partial automation of both loaders and trains, while the Kerr mine employs a conventional diesel loader and truck haulage fleet. The cost reductions result primarily from lower diesel fuel, and equipment maintenance and less ventilation power demand. Iron Cap also benefits from its proximity to the MTT and lower conveying costs. The LOM underground mining cost per total dmt milled is estimated at US$4.82 (see Table 24.19). Underground mining represents 93% of the mill feed source.

 

Table 24.19Underground Mine Operating Costs

 

Underground Mining Operating Costs Units Cost
Total Cost (LOM)  US$ million 7,559
Cost per mill feed tonne milled  US$/dmt 4.82

 

Process Operating Costs

 

The most significant process costs are reagents, grinding media, and power. Minor costs include labour, maintenance parts and assaying. The LOM average process cost, including tunnel transport costs, is US$4.31/dmt milled. With the inclusion of tailing costs, the total cost for processing is US$4.46/dmt (Table 24.20) The costs are based on 170,000 t/d plant throughput.

 

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Table 24.20Process and Tailing Management Operating Costs

 

Cost Area Cost
(US$/dmt milled)
Processing 4.07
TMF 0.15
Tunnel/Train Transport 0.24
Total 4.46
   

Reagent costs were based on test work and 2022 Q1/Q2 price quotes or in-house data. Grinding media and power costs are based on 2022 Q1/Q2 price quotes. Labour costs are based on a manpower list to reflect the 2022 PEA plant design based on an industry-focused and location-specific labour survey conducted by Seabridge.

 

Process costs are also specific to each deposit’s physical and metallurgical characteristics with each responding to differences in power and steel consumption.

 

Tunnel and material handling costs for the MTT train are included in the process operating cost. They represent US$0.24/dmt milled of the process operating cost.

 

Tailing storage and handling costs were estimated by KCB at US$0.15/dmt milled and excluded in the processing operating cost. They are derived from KCB’s current databases.

 

General and Administrative Operating Costs

 

G&A costs are costs that do not relate directly to mining or processing operating costs. G&A costs include:

 

personnel: executive management, accounting, supply chain and logistics, human resources, external affairs functions, and other G&A departments

 

expenses: including insurance, off site offices, administrative supplies, medical services, legal services, human resources related expenses, community and environmental programs, accommodation/camp costs, air/bus crew transportation, regional and property taxes, and external assay/testing.

 

The G&A costs were estimated based on the 2022 PEA head count and factored costs from previous studies. G&A costs average US$1.39/dmt milled over the LOM.

 

Site Services Operating Costs

 

Site services operating cost are estimated at US$0.50/dmt milled including the following elements:

 

personnel – general site services human power

 

site mobile equipment and light vehicle operations

 

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portable water and waste management

 

general maintenance including yards, roads, fences, and building maintenance

 

off-site operation expense

 

building heating

 

power supply to surface facilities

 

avalanche control.

 

Water Management Operating Costs

 

Water management costs are estimated at US$0.68/dmt milled and were derived from databases and current studies and adjusted to reflect the 2022 PEA requirements and to reflect current commodity pricing. Table 24.21 shows the average LOM operating costs for the HDS treatment plant, the selenium water treatment plant, and the water pumping system.

 

Table 24.21Water Management Costs

 

Description US$/t
HDS Water Treatment 0.37
Se Water Treatment 0.26
Water Pumping 0.05
Total 0.68

 

24.22Economic Analysis

 

The results of the economic analysis in the 2022 PEA represents forward-looking information that is subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here. Forward-looking statements in the 2022 PEA section of this report include, but are not limited to, timing and amount of future cash flows from mining operations, forecast production rates and amounts of copper, gold, silver, and molybdenum produced from the KSM mining operation, estimation of the Mineral Resources and the realization of the Mineral Resource estimates within the 2022 PEA mine plans, the time required to develop the mine based on the 2022 PEA mine design, statements with respect to future price of copper, gold, silver, and molybdenum, currency exchange rate between the US dollars and Canadian dollars, assumptions regarding mine dilution and losses, the expected grade of the material delivered to the mill, metallurgical recovery rates, initial capital and sustaining capital costs, as well as mine closure costs and reclamation, timing and conditions of permits required to initiate mine construction, maintain mining activities, and mine closure, and assumptions regarding geotechnical and hydrogeological factors.

 

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The reader is cautioned that the actual results of mining operations may vary from what is forecast. Risks to forward-looking information include, but are not limited to, unexpected variations in grade or geological continuity, as well as geotechnical and hydrogeological assumptions that are used in the mine designs. There could be seismic or water management events during the construction, operations, closure, and post-closure periods, that could affect predicted mine production, timing of the production, costs of future production, capital expenditures, future operating costs, permitting timelines, potential delays in the issuance of permits, or changes to existing permits, as well as requirements for additional capital. The plant, equipment or metallurgical or mining processes may fail to operate as anticipated. There may be changes to government regulation of mining operations, environmental issues, permitting requirements, and social risks, or unrecognized environmental, permitting and social risks, closure costs and closure requirements, unanticipated reclamation expenses, title disputes or claims and limitations on insurance coverage.

 

A portion of the Mineral Resources in the mine plans, production schedules, and cash flows include Inferred Mineral Resources, that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the 2022 PEA will be realized. Due to the conceptual nature of the 2022 PEA, none of the Mineral Resources in the 2022 PEA have been converted to Mineral Reserves and therefore do not have demonstrated economic viability.

 

24.22.1Methodology Used

 

The 2022 PEA has been evaluated using a discounted cash flow (DCF) analysis. Cash inflows consist of annual revenue projections for the mine. Cash outflows such as capital, including the four years of pre-production costs, operating costs, taxes, and royalties are subtracted from the inflows to arrive at the annual cash flow projections. Cash flows are taken to occur at the end of each period.

 

To reflect the time value of money, annual net cash flow (NCF) projections are discounted back to the valuation date using several discount rates. The discount rate appropriate to a specific project depends on many factors, including the type of commodity; and the level of project risks related to mine construction and operation, such as market risk, technical risk and political risk. The discounted, present values of the cash flows are summed to arrive at the NPV. Preproduction development period for the PEA was assumed to be 4 years, economic results were reported at the start of that 4 years period.

 

In addition to NPV, IRR and payback period are also calculated. The IRR is defined as the discount rate that results in an NPV equal to zero. Payback is calculated as the time require to achieve positive cumulative cash flow following first metal production.

 

Industry common practice where mineralization extends at depth beyond open pit mining is to transition the mining method from open pit to underground without pausing the mining activity to accommodate the transition. The 2022 PEA assumes the pre-development period and associated costs are exclusive from the 2022 PFS; however, an opportunity exists to improve the 2022 PEA’s business case by investing free cash flow from the tail end of the operations described in the 2022 PFS as the initial capital for the 2022 PEA.

 

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24.22.2Financial Model Parameters

 

Basis of Analysis

 

The financial analysis was based on the Mineral Resources presented in Section 14, the mine and process plan and assumptions detailed in Sections 24.16 and 24.17, respectively, the projected infrastructure requirements outlined in Section 24.18, the concentrate marketing assumptions in Section 24.19, the permitting, social and environmental regime discussions in Section 24.20, and the capital and operating cost estimates detailed in Section 24.21.

 

Metal Pricing

 

A Base Case economic evaluation was undertaken incorporating historical three-year trailing averages for metal prices as of June 20, 2022. Two alternate cases are also presented: (i) an Alternate Case that incorporates lower metal prices than used in the Base Case to demonstrate the 2022 PEA sensitivity to lower prices; and, (ii) a Recent Spot Case incorporating recent spot prices for gold, copper, silver and the US$/Cdn$ exchange rate. Metal prices for each scenario are presented in Table 24.22.

 

Royalties

 

Under the Benefits Agreement with the Nisga’a Nation and the Co-operation and Benefits Agreement with the Tahltan Nation, the project owners make an annual financial payment to each Nation once operations have commenced. The combined annual payments to these Nations are payable in two forms; payments that are a percentage of the tax payable (the “Mineral Tax”) under the Mineral Tax Act (British Columbia) (the “Mineral Tax Act”), which is a tax on net operating profit, and payments that are based on net smelter returns. Royalty payments include 60% of the gross silver royalty payable to Sprott.

 

Working Capital

 

Working capital cash outflow and inflows are included in the financial model. The calculations are based on the assumptions that accounts payable will be paid within 90 days and accounts receivable within 30 days.

 

Closure and reclamation

 

Closure costs are estimated at US$588 million.

 

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Taxes

 

The post-tax financial estimates consider all applicable Canadian Federal and BC Provincial taxes including:

 

Canadian Federal Income Tax

 

BC Provincial Tax

 

BC Mineral Tax

 

Provincial Sales Tax

 

Canadian Federal and BC Provincial Income Tax Regime

 

The federal and BC provincial corporate income taxes are calculated using the currently enacted rates of 15% for federal and 12% for BC. Over the LOM, approximately US$3,969 million in federal and US$3,175 in provincial corporate income taxes are paid. For both federal and provincial income tax purposes, capital expenditures are accumulated in tax pools that can be deducted against mine income at different prescribed rates, depending on the type of capital expenditures.

 

BC Mineral Tax Regime

 

The BC Mineral Tax regime is a two-tier tax regime, with a 2% tax and a 13% tax. Over the LOM, approximately US$4,159 in BC Mineral Tax is paid.

 

The 2% tax is assessed on “net current proceeds”, which is defined as gross revenue from the mine less mine operating expenditures. The 13% tax is assessed on “net revenue”, which is defined as gross revenue from the mine less any accumulated expenditures.

 

BC Mineral Tax is deductible for federal and provincial income tax purposes.

 

Provincial Sales Tax

 

Provincial sales tax (PST) is a retail sales tax that applies when a taxable good or service is purchased, acquired or brought into B.C., unless a specific exemption applies. The proportion of goods and services within the capital costs that are non-exempt from PST was determined to be 18% in previous studies. The PST retail sales tax is 7%. For estimating PST that applies to capital, the 7% tax was applied to 18.5% of the annual capital and sustaining capital estimates within the financial model. The LOM PST for capital and sustaining capital costs was US$19 million and US$165 million, respectively.

 

Financing

 

The model does not include any costs associated with financing.

 

Inflation

 

There is no adjustment for forward inflation in the financial model; all cash flows are based on 2022 dollars.

 

Financial Results

 

Table 24.22 summarizes the financial results. The post-tax NPV at a 5% discount rate over the estimated mine life is US$5.8 billion. The post-tax IRR is 18.9%. The post-tax payback of the initial capital investment is estimated to occur in 6.2 years after the start of production.

 

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The average life of mine operating cost per ounce of gold recovered is a negative US$-343 and US$460 on a byproduct and a coproduct basis, respectively; and the average life of mine operating cost per pound of copper recovered is US$0.38 and US$0.93 on a byproduct and a coproduct basis, respectively. The total cost in Table 24.22 is calculated similar to the operating cost but includes US$14.8 billion in capital and closure costs.

 

Caution Statement

 

The PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the PEA will be realized. Mineral Resources in the PEA mine plan are not Mineral Reserves and do not have demonstrated economic viability.

 

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Table 24.22 2022 PEA Financial Analysis Summary

 

Financial Results 2022 PEA Base Case
3yr Avg.
2022 PEA Alternative
Case

2022 PEA

Recent Spot
Upside Case

Metal Prices:      
Gold (US$/ounce) 1,742.00 1,500.00 1,850.00
Copper (US$/pound) 3.53 3.00 4.25
Silver (US$/ounce) 21.90 20.00 22.00
Molybdenum (US$/pound) 18.00 18.00 18.00
Exchange Rate: (US$/Cdn$) 0.77 0.77 0.77
Cost Summary:      
Op cost (US$)/ Au oz recovered byproduct -1,392 -905 -2,060
Total cost (US$)/ Au oz recovered byproduct -343 144 -1,011
Op cost (US$)/ Au oz recovered coproduct 460 463 423
Total cost (US$)/ Au oz recovered coproduct 1,053 1,044 984
Op cost (US$)/ lb Cu recovered net of byproduct 0.38 0.59 0.32
Total cost (US$)/ lb Cu recovered net of byproduct 1.44 1.64 1.38
Op cost (US$)/ lb Cu recovered coproduct 0.93 0.93 0.97
Total cost (US$)/ lb Cu recovered coproduct 2.13 2.09 2.26
Initial Capital (US$ Billion) 1.5 1.5 1.5
Sustaining Capital (US$ Billion) 12.75 12.75 12.75
Unit Operating Cost On-site (US$/t milled)  11.98 11.98 11.98
Pre-Tax Results:      
Net Cash Flow (US$ Billion) 29.8 19.4 40.9
NPV @ 5% Discount Rate (US$ Billion) 9.7 5.8 13.9
Internal Rate of Return (%) 24.0 17.4 30.4
Payback Period (Years) 4.7 7.5 3.9
Post-Tax Results:      
Net Cash Flow (US$ Billion) 18.5 11.9 25.6
NPV @ 5% Discount Rate (US$ Billion) 5.8 3.3 8.4
Internal Rate of Return (%) 18.9 13.5 24.0
Payback Period (Years) 6.2 8.7 4.4

Note: *June 20, 2022 3-year trailing average.

 

Sensitivity to metal prices, operating costs, capital costs, and exchange rate were analyzed for the 2022 PEA Base Case post-tax NPV 5% and IRR. The sensitivities are shown graphically in Figure 24.22 and Figure 24.23.

 

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Figure 24.22 Post-Tax NPV 5% Sensitivity Analysis

 

Source: Tetra Tech (2022)  

 

Figure 24.23 Post-Tax IRR Sensitivity Analysis

 

Source: Tetra Tech (2022) 

 

The 2022 PEA financials are more sensitive to changes in copper price and exchange rate than changes in capital costs and operating costs.

 

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25.0Interpretations and Conclusions

 

25.1Introduction

 

The KSM Property involves the development of major gold-copper-silver-molybdenum deposits located in northwest BC. KSM includes five major mineralized zones, identified as the Mitchell, East Mitchell, Kerr, Sulphurets, and Iron Cap deposits.

 

KSM has received environmental assessment approvals and early-stage construction permits. Early construction works currently underway include the establishment of access roads, construction camps, and power infrastructure.

 

In this report, this “KSM Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report” (the Report), the KSM mine development has been evaluated as a PFS (2022 PFS) using open pit mining of the Mitchell, East Mitchell, and Sulphurets deposits. A PEA (2022 PEA) is a standalone mine plan that has been undertaken to evaluate a potential future expansion of the KSM mine to the Iron Cap and Kerr deposits after the 2022 PFS mine plan has been completed. The Mineral Resources included in the PEA mine plan are only those Mineral Resources that were not included in the PFS mine plan.

 

The results of the 2022 PFS represent a viable option for developing the Property, with the PEA assessing a post-PFS mine life extension option at a conceptual level.

 

The 2022 PFS maintains the mine development scope as a large-tonnage open pit operation, with the annual average maximum nominal mill throughput increasing from initial 130,000 t/d to 195,000 t/d by Year 3.

 

The flotation and leaching plants will be capable of producing a copper/gold/silver concentrate for transport by truck to the nearby deep-water seaport at Stewart, BC. A gold-silver doré, and a separate molybdenum concentrate, will also be produced at the processing facility.

 

25.22022 Prefeasibility Study Conclusions

 

25.2.1Exploration and Mineral Resources

 

Significant drilling and other exploration activities have been conducted in the KSM district since the 1960s. A number of major mining companies completed various exploration programs prior to Seabridge’s entry into the district in 2000. Since then, Seabridge has drilled the Kerr, Sulphurets, Mitchell, East Mitchell, and Iron Cap deposits, testing both the limits and geometry of the extensively altered and mineralized systems that appear to be centered on hypabyssal, early-Jurassic intrusions that are located adjacent to regional thrust faults. The East Mitchell (formerly named Snowfield) deposit, acquired by Seabridge in 2021, is also included in the 2022 Mineral Resources. Seabridge’s geological staff developed geological models for each of the mineralized systems. Those geological models were used to create grade models that were used to tabulate Mineral Resources for the KSM Property. The KSM Mineral Resources are constrained by conceptual mining shapes based on their assumed mining methods. The KSM Mineral Resources were prepared in accordance with CIM Definition Standards (2014) and CIM Estimation of Mineral Resource and Mineral Reserves Best Practice Guidelines (2019).

 

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25.2.22022 PFS Mineral Reserves

 

All of the KSM Mineral Reserves were converted from Measured and Indicated Mineral Resources described in Section 14.0 by applying modifying factors within the 2022 PFS. Mineral Reserves are based on NSR values that were calculated in the block model using US$1,300.00/oz of gold, US$3.00/lb of copper, US$20.00/oz of silver, US$9.70/lb of molybdenum, and a foreign exchange rate of US$0.79 per Cdn$1.00, with varying process recoveries for the different mining areas, and applicable off-site charges. The NSR values have been used as a dynamic cut-off for defining ore and waste in the open pit, with a minimum NSR of Cdn$11.00/t. Mining loss and dilution parameters for the open pits mine plan are applied as described in Section 15.0. The estimated Proven and Probable Mineral Reserves as of May 26, 2022, are 47.3 Moz of gold and 7.32 Blb of copper (2.29 Bt at an average grade of 0.64 g/t gold, 0.14% copper, 2.2 g/t silver and 76 ppm molybdenum per t).

 

The relevant geotechnical, hydrology, and hydro-geology components of the mine design have been provided by competent professional consulting firms with a long history on the KSM property. These supporting studies have been used in establishing the mine design parameters and supporting facilities designs for the 2022 PFS mine plan.

 

The methods used in this estimate are in accordance with CIM Definition Standards, with reasonable engineering practices for a PFS-level study, and economic estimates based on the technical and economic parameters stated in this Report.

 

25.2.3Mining Methods

 

The open pit mine plan in the 2022 PFS establishes the economic mining limits of the Mitchell, East Mitchell, and Sulphurets Mineral Resource areas using large-tonnage mining methods capable of providing mill feed at a nominal rate of 195,000 t/d after early production years ramp up. The LOM plan accommodates the local adverse conditions comprising snow, cold, remoteness, and steep terrain. Waste and water management designs are incorporated into the mine plan, as specified in the current site plans and as reviewed and approved in the Application/EIS (Rescan 2013) review process completed in 2014.

 

The chosen mine equipment is well known and suitable for the expected operating conditions, and the productivity assumptions are reasonable and achievable. The resultant unit mining costs are comparable when benchmarked against other similar operating mines when considering the site operating conditions. Given the stated design parameters and assumptions, the open pit mine plan will achieve the forecast production schedule and the annual levels and costs within the expected range of accuracy of the 2022 PFS estimate.

 

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25.2.4Recovery Methods

 

Several wide-ranging and extensive metallurgical test programs have been conducted since 2007 to assess the metallurgical responses of the mineral samples from the KSM deposits, especially the samples from the Mitchell deposit. The test results indicate that the mineral samples from the Mitchell, East Mitchell, and Sulphurets mineralized deposits proposed for the mine production plan are amenable to the flotation-cyanidation combined process.

 

The 2022 PFS processing plant is designed based on the flowsheet developed from the testwork results. The proposed flotation process is projected to produce an average copper-gold concentrate containing approximately 24% copper with approximately 25% or higher concentrate grades in the initial years. Copper and gold flotation recoveries will vary with changes in head grade and mineralogy. The LOM average copper and gold recoveries to the concentrate and doré are projected to be 80.2% and 71.7%, respectively, for the LOM mill feed containing 0.64 g/t gold and 0.14% copper. As projected from the testwork, the cyanidation circuit will increase the overall gold recovery to a range of 60% to 79%, depending on gold and copper head grades. Silver recovery from the flotation and leaching circuits is expected to be 61.1% on average. A separation flotation circuit will recover molybdenite from the copper-gold-molybdenum bulk concentrate when higher-grade molybdenite mineralization is processed.

 

The process flowsheet proposed for KSM is conventional and has been widely used in processing porphyry copper-gold ores. The equipment type and sizing selected for KSM are modern and large equipment used in other mining operations or projects.

 

In general, the copper and molybdenum concentrates produced are anticipated to be acceptable by most of the copper and molybdenum smelters. On average, the impurity contents in the copper and molybdenum concentrates should be lower than the penalty thresholds set by most of the smelters, although it is anticipated that the arsenic and antimony may exceed the penalty thresholds set up by some of smelters in some short periods.

 

25.2.52022 PFS Project Infrastructure

 

MTT Transportation System

 

In the 2022 PFS, transportation of ore, freight, and personnel through the MTT will be achievable with a twin tunnel automated train transport system, at an average rate of 195,000 t/d and a peak capacity of 12,000 t/h after an initial three-year ramp up. The required deliveries of freight, including bulk transport of fuel and lime, mine site consumables, and personnel movement for periodic crew changes and daily requirements between the Mitchell OPC and Treaty OPC are also achievable using the tunnel and rail infrastructure with appropriate traffic control management systems.

 

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The system is scalable and flexible, as it has the ability to add or remove train sets to meet higher or lower throughput requirements and components can be taken out of operation for maintenance without compromising total system operation. The twin tunnel also allows segments to be isolated for maintenance while still meeting ongoing production requirements.

 

This scale of underground tunnel transport via trains has proven effective in other mines globally, utilizing equipment and infrastructure specified within the 2022 PFS. The planned train system operations can be reasonably accomplished at the 2022 PFS’s estimated productivities and costs.

 

Tunnelling

 

For the 2022 PFS, the conventional drill and blast methodology for excavating infrastructure and water tunnels is a valid basis for the scheduling and costing of the long tunnels required in this PFS. Using a twinned tunnel for the MTT provides advantages in construction with three sets of advancing twin headings, enabling the use of one tunnel at each heading as a fresh airway, and the other tunnel as a return airway, which has a significant impact on ventilation and advance rates for long tunnels. Other infrastructure and water tunnels, which are shorter, will be excavated with a single advancing heading from each portal, a well-proven method in the mining industry.

 

Advance rates and excavation costs for the MTT have been determined from contractor estimates and engineering simulation accounting for cycle times varying ground conditions. The contractor-developed advance rates and costs have been adapted for the other tunnel excavations. The contractor’s estimates have also been benchmarked, indicating the estimates are within the accuracy of the 2022 PFS.

 

Infrastructure Dams

 

Tailings Management Facility

 

In 2022 PFS, the TMF will be a conventional tailings storage facility with three main cells, using initial starter dams with ongoing centreline dam raises using cycloned sand. The TMF layout and sequencing allows for staged construction of the facility, which reduces both start-up and operating costs, while managing geotechnical and environmental risks. Using a three-cell system provides advantages for storage of CIL tailings, and allows progressive, staged development followed by closure of the facility to reduce environmental and closure risks. The cell system reduces TMF discharge of excess water and also allows for progressive development of water management facilities.

 

A Best Available Tailings Technology (BATT) assessment (KCB, 2016a), shows that the current TMF design is appropriate to meet environmental, operability, and geotechnical criteria. Filtered tailings (Dry stack) options were either not the most appropriate or did not meet these criteria.

 

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Key conclusions include:

 

TMF starter dams safely store 18 to 24 months of tailings under the range of start-up assumptions assessed as established by the TMF design criteria. The storage plan allows for operational flexibility during start up, with a minimum of one winter season and potentially two winter seasons accommodated

 

TMF dam designs (from starter dams through closure) are stable under static and seismic loading, as designed, and these designs meet all applicable regulatory criteria.

 

Water Storage Facility

 

The WSF provides environmental containment of runoff water for the mine site. The facility includes a rock fill-asphalt core WSD to retain contact water collected from the mine site for treatment at the HDS WTP. The facility is capable of storing inflows from extreme events and is designed to minimize seepage. The WSD will be built to full height before start-up to provide approximately 50 Mm3 of storage including flood storage from a 200-year wet year. Key conclusions include:

 

The rock fill-asphalt core WSD design is confirmed to be the preferred structure type to meet the KSM environmental, durability, and geotechnical design criteria.

 

A value engineering study (KCB, 2012e) showed that this design has both low-seepage rates, as well as constructability and cost advantages over other types of dam structures analyzed for this location and purpose

 

25.2.62022 PFS Economic Analysis

 

Tetra Tech prepared an economic evaluation for the 2022 PFS based on a pre-tax financial model. The tax component of the model was prepared and reviewed by the tax consultant (please see Section 22.0 for further details). Based on this tax analysis, Tetra Tech prepared the post-tax economic evaluation of the 2022 PFS.

 

For the 33-year LOM and 2.29 Bt Mineral Reserve, Table 25.1 summarizes the 2022 PFS economic analysis results.

 

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Table 25.1Summary of Major 2022 PFS Pre- and Post-Tax Results by Metal Price Scenario

 

  Unit 2022 PFS
Base
Case
2022 PFS
Spot
Case
2022 PFS
Alternate
Case
Gold US$/oz 1,742.00 1,850.00 1,500.00
Copper US$/lb 3.53 4.25 3.00
Silver US$/oz 21.90 22.00 20.00
Molybdenum US$/lb 18.00 18.00 18.00
Exchange Rate US$:Cdn$ 0.77 0.77 0.77
Pre-Tax Results
Undiscounted NCF US$ million 38,636 46,070 27,854
NPV (at 3%) US$ million 20,210 24,357 14,210
NPV (at 5%) US$ million 13,454 16,403 9,194
NPV (at 8%) US$ million 7,420 9,294 4,717
IRR % 20.1 22.4 16.5
Payback years 3.4 3.1 4.1
Cash Cost/oz Au US$/oz 275 164 351
Total Cost/oz Au US$/oz 601 490 677
Post-tax Results
Corporate Tax (Federal) US$ million 5,152 6,118 3,754
Corporate Tax (Provincial) US$ million 4,122 4,894 3,003
BC Mineral Tax US$ million 5,429 6,427 3,963
Total Taxes US$ million 14,703 17,440 10,721
Undiscounted NCF US$ million 23,933 28,630 17,133
NPV (at 3%) US$ million 12,264 14,889 8,467
NPV (at 5%) US$ million 7,944 9,814 5,238
NPV (at 8%) US$ million 4,061 5,254 2,332
IRR % 16.1 18.0 13.1
Payback years 3.7 3.4 4.3
Notes:1) Operating and total cost per ounce of gold are after copper, silver, and molybdenum credits.
2)Total cost per ounce includes all start-up capital, sustaining capital, and reclamation/closure costs.
3)Results include consideration of Royalties and Impact Benefit Agreements.
4)The post-tax results include the BC Mineral Tax and provincial and federal corporate taxes.
5)Sums may not add due to rounding.

 

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25.32022 Preliminary Economic Assessment Conclusions

 

The 2022 PEA has been undertaken to evaluate a potential future expansion of the KSM mine to the Iron Cap and Kerr deposits after the completion of the PFS plan. None of the Mineral Resources incorporated into the 2022 PEA evaluation have been used in the 2022 PFS.

 

The 2022 PEA is primarily an underground block cave mining operation supplemented with a small open pit and is planned to operate for 39 years with a peak mill feed production of 170,000 t/d.

 

25.3.1Mining Methods

 

The 2022 PEA mine design is a combined open pit/underground block cave mining operation. The mine production plan starts as lower-cost open pit mining at Upper Kerr, using conventional large-scale equipment while developing the Iron Cap block cave mine. After the Iron Cap mine has been developed to begin feeding ore to the mill, development of the first lift of the Lower Kerr block cave will commence.

 

The Iron Cap and Kerr underground mines will be the main source of mill feed, contributing approximately 93% of the total plant feed over the LOM, supplemented by the Kerr open pit in the mine production plan.

 

Open Pit Mining

 

The 2022 PEA open pit mining is designed as a conventional truck-shovel operation. Kerr open pit has been designed to supplement block cave mill feed during the ramp up of the block cave production.

 

There is one year of pre-stripping before production commences. Open pit production will occur from PEA Years 1 to 5 of the PEA mine plan using an already established KSM open pit mining fleet to supplement Iron Cap mill feed ramp up. Open pit production is completed before Kerr underground production begins.

 

Underground Mining

 

The 2022 PEA underground block caving mine designs for Iron Cap and Kerr are based on modeling using GEOVIA’s Footprint Finder software. Iron Cap has been designed for using battery-electric loaders and electric trains for material handling, and employs both tele-operation and automation for these units. This enables a reduction to the number of primary ventilation intake and exhaust drifts due to less ventilation demand and operating cost savings from lower diesel fuel consumption, labour, and equipment maintenance costs. The Kerr block cave mine is designed as a conventional diesel equipment mine, presenting a future opportunity to electrify.

 

The first underground mill feed production from Iron Cap and Kerr comes in PEA Years 1 and 7 of the PEA mine plan, respectively. Iron Cap is estimated to have a production ramp-up period of 6 years, steady state production at 32.9 Mt/a for 17 years, and then ramp-down production for another 6 years. Kerr is estimated to have a production ramp-up period of 5 years and steady state production at 29.2 Mt/a for 20 total years, with a 7-year production dip during years where the operation transitions from the first to second lift. Underground production ends first at Iron Cap in PEA Year 29 and finally at Kerr in PEA Year 39.

 

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25.3.2Recovery Methods

 

Based on the available information, including preliminary mine plan and results of the metallurgical test work performed on various samples providing a reasonable indication of the mineralogical characteristics of the materials, the process flowsheet developed for the KSM mineralization is considered appropriate for the 2022 PEA.

 

Extensive metallurgical test programs have been carried out to assess the recoverability of copper, gold, silver, and molybdenum values using the flowsheet developed for the KSM mineralized material. The results of the test programs indicate that the mineral samples from the Iron Cap and Kerr deposits are amenable to the flotation process method developed for the KSM mineralization.

 

The copper-gold concentrates from the mineralization are expected to be saleable with no significant penalty elements. Cyanidation tests showed that the bulk cleaner flotation tailings and pyrite concentrate were possible to recover gold and silver by cyanide leaching, however, preliminary trade-off studies indicate that cyanidation of the Iron Cap and Kerr gold-bearing sulphide materials may not be economic due to ineffective gold and silver absorption onto activated carbon grains caused by elevated copper/gold ratios. Additional test work and economic assessments are required to validate this observation. The metallurgical test work also determined the consumption of reagents and grindability which have been incorporated into the operating cost estimates.

 

The metallurgical performance parameters for copper, gold, silver, and molybdenum minerals are projected based on the metallurgical test results obtained from various test programs that are summarized in the metallurgical test work review section.

 

Detailed characterization and metallurgical test work on Iron Cap and Kerr samples are presented in the pertinent test work reports listed in Section 27.0.

 

25.3.32022 PEA Project Infrastructure

 

MTT Transport System

 

Mill feed for the PEA will be delivered from the mining operations to the Mitchell OPC then will be transported to the Treaty OPC via the twin tunnels and automated train system. This facility will also provide access for personnel transport, fuel, supplies, and the power supply to the Mitchell area.

 

Tailings and Water Management

 

TMF management philosophy and criteria for the 2022 PEA are generally consistent with the TMF designs presented in the approved EA. The 2022 PEA has a lower peak throughput, 170,000 t/d, which is less than the process plant capacity, so tailings storage would account for this modified placement rate and maintain requisite supernatant pond operational volume and freeboard. A high-level allowance for expanded total tailings production is included in the costing.

 

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The water and waste management structure designs in the 2022 PEA are based on those described in the approved EA. The approximate length of existing single axis water diversion tunnels in Mitchell Valley at PEA Year 1 is approximately 18 km and includes the MDT, MTDT, NPWDA and MVDT.

 

The 2022 PEA rock storage requirements are significantly less than in the PFS as there is only one deposit mined by open pit methods producing mine waste (i.e., Kerr). The 2022 PEA waste management plan is compatible with the approved EA by consolidating waste rock into a single facility, including minor waste rock from initial underground development, into Mitchell Valley as pit backfill which is contiguous with the Mitchell RSF.

 

The overall 2022 PEA operational water management and water treatment designs are similar to those described in the approved EA, and no adjustments were required since the three areas to be mined in the 2022 PEA are inside disturbance areas addressed in the approved EA.

 

The WSD crest elevation is kept at the same elevation for the 2022 PEA and WSD operations would be consistent with those describe for the approved EA.

 

25.3.42022 PEA Economic Analysis

 

A Base Case economic evaluation was undertaken incorporating historical three-year trailing averages for metal prices as of June 20, 2022. This approach is used because it is consistent with the 2022 PFS Base Case. Two alternate cases are also presented:

 

An Alternate Case that incorporates lower metal prices than used in the Base Case to demonstrate 2022 PEA’s sensitivity to lower prices; and,

 

A Recent Spot Case incorporating recent spot prices for gold, copper, silver, and the US$/Cdn$ exchange rate.

 

In the 2022 PEA, it is assumed that the pre-production development period will be four years long. The base date of the economic results presented as follows is at the beginning of this period. The estimated 2022 PEA pre-tax and post-tax economic results are presented in Table 25.2.

 

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Table 25.22022 PEA Financial Analysis Summary

 

Financial Results 2022 PEA
Base Case
3-yr Avg.
2022 PEA
Alternative
Case
2022 PEA
Recent Spot
Upside Case
Metal Prices:      
Gold (US$/oz) 1,742.00 1,500.00 1,850.00
Copper (US$/lb) 3.53 3.00 4.25
Silver (US$/oz) 21.90 20.00 22.00
Molybdenum (US$/lb) 18.00 18.00 18.00
Exchange Rate: (US$/Cdn$) 0.77 0.77 0.77
Pre-Tax Results:
Net Cash Flow (US$ Billion) 29.8 19.4 40.9
NPV @ 5% Discount Rate (US$ Billion) 9.7 5.8 13.9
Internal Rate of Return (%) 24.0 17.4 30.4
Payback Period (Years) 4.7 7.5 3.9
Post-Tax Results:
Net Cash Flow (US$ Billion) 18.5 11.9 25.6
NPV @ 5% Discount Rate (US$ Billion) 5.8 3.3 8.4
Internal Rate of Return (%) 18.9 13.5 24.0
Payback Period (Years) 6.2 8.7 4.4

 

The 2022 PEA offers a viable option for extended development of the very large underground Mineral Resources at KSM after the 2022 PFS Mineral Reserves are depleted.

 

25.42022 Prefeasibility Study Risks

 

There are risks that could affect the economic viability of the KSM mine development. Many of these risks are based on the current extent of sufficiently detailed information and engineering in specific areas. These risks can be further managed as more drilling, sampling, testing, design, and engineering are completed in the next study stage.

 

However, several risks have been reduced through activities completed by Seabridge and its team since 2012. Specifically, the permitting risk has been addressed substantially with the receipt of environmental approvals granted by both the federal and provincial governments in 2014 and associated approved extensions, and the granting of the early-stage construction permits for KSM including the MTT License of Occupation. It can be effectively stated that Seabridge earned “the social license” for KSM through the successful completion of the environmental review process with the support of nearby communities (by submission of letters of support) and Indigenous groups. On June 16, 2014, Seabridge Gold entered into a comprehensive Benefits Agreement with the Nisga’a Nation in respect of the KSM Property. On July 8, 2019, the Tahltan Nation and Seabridge Gold Inc. announced that the parties reached agreement on the terms of a Co-operation and Benefits Agreement in connection with KSM.

 

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One of the significant unknowns associated with KSM is related to the extent of available geotechnical data for the water diversion tunnels. Since several tunnels are key for the development of the Mine Site, potential delays in construction could occur if unforeseen rock mass conditions or groundwater inflows are encountered that cause durations longer than those anticipated in the preliminary construction schedule. The MTT tunnel geotechnical drilling and geotechnical characterization have been completed. Geotechnical drilling along the mine water tunnel routes is currently underway, and it is anticipated that this mine water tunnel geotechnical data gap will be reduced at the completion of the ongoing 2022 site investigation program.

 

A significant risk common to all aspects of the KSM mine development is the availability of labour and experienced management and supervision. There are currently open pit and underground mines operating in the area surrounding KSM, thus generating a locally experienced labour pool; however, the shortage of experienced labour is currently a global mining industry problem. This risk will need to be addressed by KSM with a robust recruitment and training program.

 

These risks are common to most mining projects, many of which can be mitigated with adequate engineering, planning, and proactive management. Some external risks, such as metal prices, exchange rates, inflation, and government legislation, are beyond the control of the developer and operator and are difficult to anticipate and mitigate, although in some instances measures for risk reduction have already been included in conservative design such as in selection of economic mining limits. Risk reduction measures are also included in tunnel and access road design, TMF design, and through the inclusion of early scheduling for items potentially subject to risk of delay. The means to address risk for a project the size of KSM moving forward is to establish a formal risk management program during advanced study phases that continues through development and into mine operation. The KSM project team will systematically review risks and opportunities during project development and construction, and take appropriate action to minimize the impact on overall costs and scheduling.

 

Table 25.3PFS Risks and Mitigation Measures

 

Risks Mitigation Measures
Open Pit Mining
Slope stability, safety, and production delays. Throughout the planning, design, and engineering of the open pits, the KSM project team has identified hazards and developed mitigation plans to reduce and manage risks related to the open pit slopes.
Slope deformation or rock falls from unstable slopes requiring restricted access to the pit, slope rehabilitation, and lost or reduced mine production. Appropriate geotechnical design and assessment prior to mining, a comprehensive slope monitoring and management plan during mining, the addition of extra-wide benches at regular 150 m intervals in the pit slope configurations, and standard operating procedures that include good wall control practices and mine operations planning.
  table continues…

 

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Risks Mitigation Measures
Snow avalanches at the pit crests resulting in restricted access to the pit. An avalanche management plan coordinated with mine operations.
Visibility or weather shutdowns resulting in reduced productivity in the open pits. Snow handling crews and equipment, planned lost days, and the use of ore stockpiles.
Slope depressurization to reduce high pore water pressures in the pit slopes that might result in pit slope instability. An extensive slope depressurization plan that includes vertical wells, horizontal drains, and dewatering adits as a multi-layer system to achieve the design depressurization targets.
Surface water reporting to the pits due to failure or inundation of water management structures. Excessive surface water into the pits may result in localized slope failures, reduced mine productivity, or increased pumping costs. Adjusting the mine plan to limit the exposure of haul ramps to potential erosion from surface water sources and adequate geotechnical design of the water management infrastructures.
Large open pits mined in areas of mountainous terrain with high precipitation. The height of the highwalls also needs consideration. The mining and geotechnical teams have drawn on experience from design work in similar terrain and operations in these conditions and applied it to the designs in this study to reduce the risk through design and operating practices and procedures integrated into the plan.
Tunnels
Surface routes for ore transport and surface water diversion are more susceptible to conflicts with other surface infrastructure and ongoing mine operations, and also present risks from climate and geohazards. Tunnels provide direct routes for ore transport from the mining areas to the Treaty OPC and divert surface water around mining areas and facilities. Tunnels were assessed as having lower operational risks than alternative surface routes.
Uncertainty in permitting risk. The permit covering the MTT route was secured from the BC Government in September 2014.
Poor ground conditions caused by unforeseen faults, areas of weaker than anticipated rock, and/or higher than assumed differential stresses, and unforeseen high water inflow rates that can slow down tunnel advance rate and increase costs.

Reasonable assumptions made from preliminary-level investigations that include geological mapping of tunnel routes, geophysical surveys, drilling, and hydrogeological/geotechnical sampling and testing.

  Significant drilling along the MTT route with geotechnical characterization has been completed.
  Geotechnical drilling and characterization along the mine water tunnel alignments are currently underway.
  Use of probe and pilot/cover drilling to assist in characterizing ground conditions ahead of tunnel advance.

 table continues…

 

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Risks Mitigation Measures
Impact of construction schedule delays. The use of twinned tunnels to provide opportunities to advance alternate headings in problem areas, drilling and geophysical testing from surface.
  The use of probe and pilot/cover drilling in areas of difficult ground conditions.
  Obtaining the now-in-place permitting, allowing early tunnelling starts.
  Helicopter support prior to access road completion.
  Use of a faster to install primary MDX™ bolts followed by secondary bolts, instead of base case resin bar, will increase tunnel advance rates.
  Starting the MTT tunnel excavation as part of the early works significantly reduces the project schedule risk, particularly with early tunnel advance along the Mitchell to Saddle segment of the MTT.
Tailings Management Facility
TMF foundation conditions may vary adversely from PFS assumptions, requiring additional measures that result in an increase in TMF construction costs. TMF foundation geotechnical and hydrological investigations to collect in-situ and laboratory data started in 2022, and are planned for 2023 and beyond, to support a FS and to update the TMF engineering analyses.
Source of suitable TMF materials of construction may be further from the current assumptions resulting in an increase in TMF construction cost. KSM to complete investigations of quarries and borrow materials within the TMF impact area of the TMF and other site sources to ascertain the geotechnical and geochemical properties for use in dam construction, and to identify nearest locations of most suitable material for each TMF component.
TMF surface water inflow estimates may differ from those applied in the EA that could result in operational performance variance if not forecasted in engineering designs appropriately for surface water management. Update estimates of meteorology and hydrology using site data collected since the EA, applying current methodologies for greater reliability for estimating large storm event volumes for retention or water management structures (e.g., diversion channels, spillways and tunnels.)
The potential for TMF cyclones to not produce enough sands to meet TMF construction rate of rise requirements could limit process throughput, particularly in the early years of the mine life. Cyclone testing and geotechnical analysis to be carried out on process tailings samples.
  Identify suitable contingency borrow sources to make up any potential unforeseen cycloned sands shortage.
  Identify opportunities for advancing the construction of dam components to utilize cycloned sand surplus or by other materials of construction.

table continues…

 

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Risks Mitigation Measures
Mine Site Water Management
Characterizing flow and quality of influent to WSD and HDS WTP system is key to successful water management in the mine area. Higher than anticipated flows or water chemistry could result in additional capital and operating cost to expand treatment facilities. Updating estimates of meteorology and hydrology through using site data collected since the EA, applying current methodologies for estimating runoffs that are used on similar large watersheds in the coastal region.
  Constructing a site-wide water balance tool with technologically-relevant software and building in flexibility for evaluating a wide range of scenarios to assess impacts on water management system design and operation.
  Continued ABA/ML testing of proposed mine waste products to further refine geochemical signatures of mine waste constituents, how they react over time, and what resultant RSF effluent is predicted to be as the primary influent to the water retention and treatment system.
  Evaluation of improved water treatment technologies and management approach geared toward specific streams of influent from various mine site sources (dewatering wells, tunnel water, surface water, and RSF effluent), to yield improved discharge water quality.
  Evaluate waste rock engineering opportunities to minimize oxidation of waste.
WSD geotechnical foundation conditions may vary from current design assumptions. There is a potential for voids in faulted calcareous sediments in the WSD footprint. If foundations conditions are poorer than current design assumptions, there could be an adverse impact in capital and development schedule. Current site investigation is collecting the necessary geotechnical data to confirm the geotechnical characterization of the WSD foundation.
  Poor rock conditions may require excavation of poor rock zones to a suitable depth and potentially additional grouting.
Metallurgical Performance and Process
Lower than expected metal recoveries and copper grade of the concentrate produced. Additional testwork to better characterize metallurgical response of mineralization from the different ore sources.
     
Lower than expected performance from the HPGR comminution circuit due to higher than expected moisture and/or clay content can result in a reduction of mill production. Further process condition and flowsheet optimization would improve process design to accommodate potential variations in metallurgical performance.

 

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25.52022 Preliminary Economic Assessment Risks

 

Table 25.4PEA Risks and Mitigation Measures

 

Risks Mitigation Measures
Open Pit Mining
Slope stability, safety, and production delays. Throughout the planning, design, and engineering of the open pits, the KSM project team has identified hazards and developed mitigation plans to reduce and manage risks related to the open pit slopes.
Underground Mining

Impacts of uncertainties on meeting the production and mill feed schedules such as:

●  Design assumptions for all of these are within demonstrated industry experience and are considered achievable.

 

The rate at which the construction of the drawbells is accomplished The development advance rate during the critical pre-production period, and subsequently in the production period, have been demonstrated to be industry achievable.
The production ramp-up curve and maximum rate of draw of individual drawpoints    
The number of drawpoints available for production at any one time, and the travel distance for the LHD vehicles.    
Battery electric vehicles may not achieve production and cost estimates Assumptions are based on manufacturers input and limited industry experience due to newness of technology
Uncertainties in predicting the fragmentation of the caved rock at the drawpoints. Other block caving mines such as Palabora Mine have demonstrated that with careful planning and the availability of equipment to deal with oversized rock, drawpoint production rates comparable to those proposed at Kerr and Iron Cap are achievable.
The ore pass and grizzly systems do not achieve the planned production rate, and the material transported to the passes is coarser than expected.

Employing additional mobile breakage equipment and redesigning the undercut blasting to enhance the fragmentation during the early stages of the column draw.

  Pre-conditioning of the rock mass by hydraulic fracturing, which is expected to enhance the fragmentation of the material reporting to individual drawpoint.
Uncertainties in geotechnical characterization

Well-positioned geotechnical boreholes that have been geotechnically logged in detail and tested hydrogeologically.

  Geotechnical drill holes in the general vicinity of the proposed access development that provide adequate indications of the quality of the rock mass to minimize major uncertainties.
  Undertaking further geotechnical drilling in the future.

table continues…

 

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Risks Mitigation Measures
Abutment stresses and the interaction of the progressive advance of the undercut on the underlying drawpoints may cause more adverse stress conditions than currently estimated. Adopting a stress-shadowing advanced undercut approach, instead of the currently proposed post undercut approach.
Unforeseen stress conditions and the adverse impact of secondary blasting. Include ongoing rehabilitation of drawpoints and mucking drives.
The extent of surface disturbance from caving and the associated formation of the crater and more peripheral surface cracking may impact surface infrastructure. Major infrastructure that might be critically impacted can be conservatively located well beyond the potential impact zone.
Inflow of surface runoff from rainfall and snowmelt within the catchment area formed by natural slopes. A water management system does not rely on any temporary storage of water in any of the mine openings otherwise required for the operation of the mine.
  Dedicated water diversion tunnels, a gravity drainage tunnel at Iron Cap, and a dedicated storage capacity and pumping system included at Kerr.
The development of voids beneath the back of the active cave, and the associated concerns about hazardous conditions developing and air blasts occurring. Maintaining good knowledge about the cave profile as cave mining progresses by ongoing monitoring of the geometry of the caved material using techniques such as microseismic monitoring and seismic tomography.
Mud rushes due to the increased presence of fines. Monitoring of the flow of water from individual drawpoints.
  Temporary closure of certain areas that are deemed vulnerable until water flows at drawpoints decrease sufficiently.
  Adoption of remote mucking until conditions are deemed to be acceptable to return to full entry.
Metallurgical Performance and Process
Lower than expected metal recoveries and copper grade of the concentrate produced. Additional testwork to better characterize metallurgical response of mineralization from the different ore sources.
Lower than expected performance from the HPGR comminution circuit due to higher than expected moisture and/or clay content can result in a reduction of mill production. Further process condition and flowsheet optimization would improve process design to accommodate potential variations in metallurgical performance.

 table continues…

 

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Risks Mitigation Measures
Geological and Grade Continuity Uncertainty
A significant portion of the planned mill feed for the PEA is in the Inferred category. The location, extent, and quality of the Inferred Mineral Resource may be different from what has been modeled. Additional drilling from underground to improve confidence in the Mineral Resources when access is gained during development.
 

Mass mine method of block cave can tolerate a certain variation in the modeled location, extent and quality of the mineral resource.

 

 

The PEA mine plan relies on open pit mining of the upper Kerr deposit during the payback period where the majority of the Mineral Resources are in the Indicated category.

 

  There is adequate time to increase the mineral resource confidence categories prior to committing to the PEA mine development option.

 

25.6Mineral Resource Risks

 

Table 25.5Mineral Resource Risks

 

Risks Mitigation Measures
Mineral Resources
Changes in the geological model due to unrecognized faults and other geological structures. Update long term and short term mine plans with production data and mapping.
Large amount of Inferred in the parts of the deposits amenable to block cave mining. Infill drilling when underground access has been established.

 

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26.0Recommendations

 

26.1Introduction

 

It is recommended that Seabridge focus on advancing the KSM development by completing the data collection required to conduct a FS. The majority of the FS data collection summarized in Table 26.1 is geotechnical.

 

Infill drilling of the Iron Cap and Kerr deposits to upgrade classification will be considered at a future date.

 

Table 26.1Summary of Recommendations and Associated Costs

 

Program Cost (US$ million)
Pit Slope Geotech 4.2 to 5.3
Rock Storage Facilities Geotech 0.5
Metallurgical Testing 4.0 to 5.0
Water Storage Dam Geotech 1.0 to 1.5
TMF Geotech 8.0 to 10.0
Mine Water Tunnel Geotech 3.0 to 4.0
Site Infrastructure Geotech 1.0
TOTAL 21.7 to 27.3

 

26.2Data Collection to Complete A Feasibility Study

 

The following data collection is required before a FS can be completed.

 

26.2.1Pit Slope Geotech

 

The following geotechnical and hydrogeological work is recommended to support future assessments of the KSM open pit slopes:

 

complete a field program of large-scale hydrogeological testing via pumping wells. The results of these tests are required to increase the confidence in the design of the depressurization system for the open pit slopes of the Mitchell, East Mitchell ,and Sulphurets zones. A total of five pumping wells and ten monitoring wells for a total meterage of approximately 3,000 m is required

 

complete a field program of geotechnical drilling in the Sulphurets zone to support updates to the geotechnical model and slope designs. It is estimated that five geotechnical holes with a total meterage of approximately 2,500 m is required. A program of televiewer surveys of the drill holes, geotechnical logging of the core, and laboratory testing should be undertaken

 

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update the hydrogeological model for the latest geological model and mine plans. This update is required to improve the reliability of the depressurization quantity estimates

 

extend the slope design updates to the east and west walls of the Mitchell open pit. Develop a set of updated slope design recommendations for the whole zone that reflect the updates to the geotechnical model

 

update the slope design recommendations for Sulphurets open pits based on the results of the recommended geotechnical drilling.

 

The recommended work is estimated to cost US$4.2 million to US$5.3 million. This estimate includes drilling, drill pads, field support (helicopters, camp), and engineering fees. The range of costs are based on per drilled metre expenditures estimated previously.

 

26.2.2Rock Storage Facilities Geotech

 

KCB recommends that a geotechnical site investigation program be carried out for further delineation and characterization of foundation soils in Mitchell Valley, within the footprint of the Mitchell RSF, with a focus on the area of a known deposit of weak soils near the MTT muck piles. The characterization would use Standard Penetration (SPT) testing and undisturbed sampling for collection of samples for laboratory strength testing. This program will include:

 

geotechnical drillholes in the RSF toe above the WSF pond area. The holes should be advanced up to 60 m in overburden, with Shelby tube samples and SPT tests. Mobilization of a suitable rig for this program can be combined with drilling in the TMF area for borrow and foundation assessment

 

test pits to assess surface soil conditions may be used to augment drilling (for early-stage investigations, these will require helicopter-supported equipment mobilization)

 

geotechnical testing program including consolidation, triaxial, direct shear, and index testing.

 

The cost of this drilling and a subsequent geotechnical laboratory program is US$400,000.

 

A stability assessment and design review is required to assess the findings of the site investigations and to review suitability of existing designs to mitigate the presence of this known weak soil layer. The cost of the assessment and design review is approximately US$100,000.

 

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26.2.3Metallurgical Testing

 

In order to optimize process conditions and establish design-related parameters and more accurate metallurgical performance projections for the next stage of study, further metallurgical test work is recommended, especially the locked cycle flotation tests and cyanidation tests on various ore composite samples from the Mitchell, East Mitchell, and Sulphurets. Variability tests are also recommended for the Mitchell and East Mitchell deposits. Tetra Tech makes the following recommendations:

 

additional metallurgical test work and mineralogical evaluations should be conducted to further optimize process conditions and to establish design-related parameters for the next stage of study. The test work should include variability testing of samples from all deposits, especially from the Mitchell, East Mitchell, and Sulphurets deposits. The variability tests should be a part of the geo-metallurgical testing program that is recommended for further studies to better understand metallurgical responses of the mineralization in these resources. The test work should further investigate the effect of low slurry solid density on copper and gold in bulk rougher flotation. The test program will provide inputs for geological modelling development and mine plan update for better mill feed control

 

technically and economically feasible gold recovery processes from the gold-bearing products rejected from copper and gold bulk flotation of the KSM mineralization, such as first cleaner scavenger tailing and pyrite concentrate, should be further investigated. The processes should include further optimizing cyanide leach conditions, improving extracted gold and silver recovery from cyanide leach solution, and alternative gold and silver extraction processes, such as thiosulphate leaching, ultrafine regrinding treatment, and bacterial oxidation

 

further study should be conducted to optimize the proposed cyanide recovery and destruction methods. The study should include large scale tests using cyanidation+SART+AVR combined process for gold/silver/copper recovery to verify the process economics. Additional value that may contribute from associated copper due to co-dissolution during gold and silver extraction should be investigated. The test work should also focus on the optimum flowsheet development for the upper East Mitchell mineralization

 

a metallurgical laboratory test program should be performed on ore composites representing each year of the initial seven years of open pit mine production from Mitchell, East Mitchell and Sulphurets, according to the updated mining plan

 

further study into economical water treatment methods is recommended for water from the CIL pond.

 

The estimated costs of these programs range from US$4.0 million to US$5.0 million inclusive of sample shipping and short-term storage costs. Costs for sample collection (i.e., drilling) are not included, as this recommendation assumes samples would be derived from drill cores stored on site or in off-site warehouses.

 

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26.2.4Water Management

 

Water Storage Facility and Mine Site Water Management

 

KCB recommends additional hydrogeological and geotechnical site investigations in the WSD footprint area including:

 

hydrogeological/geological data gap analysis and review of additional water level and time series piezometer data to interrogate interaction of geological model assumptions with hydrogeological model parameters

 

additional geological mapping and geophysical surveying of the WSD footprint area. May include high-resolution drone LiDAR and photogrammetry to image and map detailed topography and geology of the WSD canyon area

 

review locations and requirements for the drilling of three large diameter wells for pumping tests, with associated monitoring holes

 

test pit programs and geotechnical laboratory testing program over the WSD footprint area to assess soil conditions

 

Based on drilling and geophysical quotes, the cost of these site investigation programs is US$1.0 million to US$1.5 million, including an allowance for drill pads, instrumentation, pumping test costs field support (helicopters, camp), and engineering fees.

 

26.2.5TMF Area

 

KCB recommends additional hydrogeological and geotechnical site investigations in the TMF footprint area consisting of:

 

hydrogeological data gap analysis and review of requirements for additional water level and time series piezometer data. Data gap analysis to refine requirements for locations of pumping tests or other investigations

 

data gap review of geotechnical (foundation) and geochemical (characterization of borrow and diversion excavations) data requirements; seismic surveying and heli-portable auger soil sampling to further delineate borrow resources and characterize construction materials

 

geotechnical drilling with hydrogeological testing (packer, multilevel piezometer, and data logger installations) at the North, Saddle and Southeast dams. Total of up to: 20 drillholes for foundation and borrow, 11 large diameter wells for pump testing, and 7 cone penetration tests are recommended for dam foundation assessment. Test pitting for borrow investigations may require helicopter support for equipment mobilization

 

seep mapping, overburden characterization, and additional overburden permeability testing in the CIL Residue Cell area to further inform the next design stages for determination of drain requirements beneath the Saddle Dam and CIL Cell liner.

 

Based on geophysical and drilling quotes, costs for these site investigation and laboratory testing programs is US$8.0 million to US$10.0 million including an allowance for line cutting, drill pads, instrumentation, pump testing, field support (helicopters, camp), and engineering fees.

 

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26.2.6Tunnels

 

MTT drilling and design is complete for construction, portal infrastructure investigations were complete in 2021, and MTT portal infrastructure design is currently ongoing. Drilling is currently underway on the alignment of the Mitchell Diversion Tunnel. The following recommendations are for investigating the diversion tunnel portals and the alignment of the remaining water diversion tunnel required at start-up which is the McTagg Diversion Tunnel:

 

Geotechnical

 

KCB recommends additional geological and geotechnical investigations at the portal locations where these have not been investigated already, and where feasible along the tunnel alignments in the PFS design. The recommended investigations consist of:

 

data gap analysis for tunnel design parameters including geological, geotechnical, and geochemical aspects

 

additional mapping and rock sampling to better characterize properties of lithological units at the portals and along the alignments for geotechnical and geochemical assessments

 

more detailed mapping of portal areas and targeted geophysical investigations to assess locations of potential structures (e.g., faults, contacts, or water bearing structures along tunnel alignments)

 

drilling for diversion tunnels and will primarily focus on shallower holes, with focus on portal and inlet area rock mass characteristics and sampling of lithological units. Several longer holes (up to 800 m) are recommended to test regional fault systems

 

geotechnical and geochemical laboratory testing on samples obtained from drilling and mapping programs.

 

The programs will inform selection of the optimum tunnelling method, allow refinement of tunnelling risk, better identify diversion tunnel lining requirements, portal locations, and portal development designs, and assist with determination of appropriate contingencies. The cost of these programs is estimated to range between US$3.0 million and US$4.0 million, including drilling costs, drill pads, field support (helicopters, camp), and engineering fees.

 

26.2.7Site Infrastructure Geotech

 

Geotechnical drilling should be done to support the design of plant infrastructure and other buildings on the mine site. A nominal allowance of 20 holes geotechnical holes averaging 25 m deep each is estimated to be required. The approximate cost for this is US$1.0 million.

 

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27.0References

 

1985 Fisheries Act, RSC. C. F-14. (1996a).

 

Alldrick, D.J., and Britton, J.M. (1988). Geology and mineral deposits of the Sulphurets area. British Columbia Ministry of Energy, Mines and Petroleum Resources, Open File Map, 4 p.

 

ALS Canada, Ltd., ALS Metallurgy Kamloops (2013). Preliminary Metallurgical Testing – Kerr Deep and Camp Zone KSM Project – (KM 3735). May 9, 2013.

 

ALS Canada, Ltd., ALS Metallurgy Kamloops (2014). Preliminary Metallurgical Testing – Deep Kerr and Iron Cap Zones – KSM Project - (KM 4029, Part A). May 16, 2014.

 

ALS Canada, Ltd., ALS Metallurgy Kamloops (2014). Preliminary Metallurgical Testing – Deep Kerr and Iron Cap Zones – KSM Project – (KM 4029, Part A). May 27, 2014.

 

ALS Canada, Ltd., ALS Metallurgy Kamloops (2015). Preliminary Metallurgical Testing – Deep Kerr – KSM Project – (KM 4514). August 05, 2015.

 

ALS Canada, Ltd., ALS Metallurgy Kamloops (2015). Preliminary Metallurgical Testing – Iron Cap – KSM Project – (KM 4672). August 04, 2015.

 

ALS Metallurgy (2015). Tailings Generation Test Work for KSM Deep Kerr and Mitchell Composites – KSM Project (KM 4811). September 3, 2015.

 

ALS Metallurgy (2016). Metallurgical Testing – Kerr Deep and Mitchell – KSM Project (KM 5087). September 27, 2016.

 

ALS Metallurgy (2017). Preliminary Metallurgical Testing – Deep Kerr Zones – KSM Project (KM 5063). January 18, 2017.

 

ALS Metallurgy (2017). Preliminary Metallurgical Testing – Deep Kerr Zones – KSM Project (KM 5266). April 13, 2017.

 

ALS Metallurgy (2017). Preliminary Metallurgical Testing – Iron Cap Zones – KSM Project (KM5248). April 13, 2017.

 

ALS Metallurgy (2018). Metallurgical Testing of a Mitchell Bulk Cleaner Scavenger Tailings Sample (KM 5455). June 19, 2018.

 

ALS Metallurgy (2018). Preliminary Metallurgical Testing – Iron Cap Zones – KSM Project (KM5501). April 17, 2018.

 

ALS Metallurgy (2018). Selenium Testing on Iron Cap Process Samples, Revision 2 (KM5501, Selenium). April 19, 2018.

 

ALS Metallurgy (2019). Preliminary Metallurgical Testing – Iron Cap Zones – KSM Project (KM5806). March 12, 2019

 

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ALS Metallurgy (2020). Preliminary Metallurgical Testing – Iron Cap Zones – Composites IC-2018-13 to IC-2018-18 – KSM Project (KM6004). January 8, 2020.

 

ALS Metallurgy (2022). Metallurgical Testing KSM Project –Mitchell Zones – KSM Project (KM6359). February 26, 2022.

 

ALS Metallurgy (2022). Metallurgical Testing KSM Project – Mitchell Zones – KSM Project (KM6461). May 11, 2022.

 

ALS Metallurgy (2022). Metallurgical Testing KSM Project – East Mitchell Zones – KSM Project (KM6461). July 22, 2022.

 

BQE Water (2022), Bench Scale Testing of SART and Gold Loading on Carbon using Leach Solutions Generated during KSM Metallurgical Testwork – July 26 2022

 

BC MEMPG (2009). Guide to Processing a Mine Project Application under the British Columbia Mines Act. British Columbia Ministry of Energy, Mines and Petroleum Resources, January 2009. http://www.coalwatch.ca/sites/default/files/Guide-to-Processing-A-Mine-Project-Application-Under-The-British-Columbia-Mines-Act.pdf.

 

BC MEMPR (2008). Health, Safety and Reclamation Code for Mines in British Columbia. British Columbia Ministry of Energy, Mines, and Petroleum Resources. http://www.empr.gov.bc.ca/Mining/HealthandSafety/Documents/HSRC2008.pdf (accessed November 2010).

 

BC MFLNRO, 2012, Ministry of Forests, Lands, Natural Resource Operations and Rural Development Engineering Manual.

 

BC MOE (2013). The Effluent Permitting Process under the Environmental Management Act: An Overview for Mine Project Applicants. Ministry of Environment, April 2103. http://www.env.gov.bc.ca/epd/industrial/mining/pdf/effluent_permitting_guidance_doc_mining_proponents_apr2013.pdf

 

BC MOF and BC MOE (1995). Riparian Management Area Guidebook. Government of British Columbia: Victoria, BC.

 

BGC (2012). Preliminary Assessment of Open Pit Slope Instability Due to the Mitchell Block Cave. December 24, 2012.

 

BGC (2020). KSM Project – Mitchell Zone – 2019 Geotechnical Model and M1/M2 Slope Designs – Rev1. Report issued to KSM Mining ULC, February 28, 2020.

 

BGC (2022). KSM Project – Design Basis for the KSM Open Pit Slopes. Report issued to KSM Mining ULC, August 05, 2022

 

Brenda Mines Ltd. (1989). Preliminary Metallurgical Testwork on “106” Low Grade Sample. Peachland, BC. May 1989.

 

Bridge, D. J. (1993). The deformed Early Jurassic Kerr copper-gold porphyry deposit, Sulphurets gold camp, northwestern British Columbia. Unpublished M.Sc. thesis, Vancouver, Canada, The University of British Columbia, 303 p.

 

Canadian Institute if Mining, Metallurgy and Petroleum (CIM) (2014). CIM Definition Standards for Mineral Resources and Mineral Reserves. May 10, 2014.

 

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CDA (2007). Dam Safety Guidelines. http://www.imis100ca1.ca/cda/CDA/Publications_Pages/Dam_Safety_Guidelines.aspx (accessed October 2012).

 

CIM Estimation of Mineral Resource and Mineral Reserves Best Practice Guidelines, November 29, 2019.

 

Cloos, M. (2001). Bubbling magma chambers, cupolas, and porphyry copper deposits. International Geology Review, 43(4), 285-311.

 

Contaminated Sites Regulation, BC Reg. 375/96.

 

DFO (1986). Policy for the Management of Fish Habitat. DFO/4486. Fish Habitat Management Branch, Department of Fisheries and Oceans Canada: n.p.

 

Environmental Management Act, SBC. C. 53.

 

Febbo, G. E., Kennedy, L. A., Savell, M., Creaser, R. A., and Friedman, R. M. (2015). Geology of the Mitchell Au-Cu-Ag-Mo porphyry deposit, northwestern British Columbia, Canada. Geological Fieldwork 2014, British Columbia Ministry of Energy and Mines, British Columbia Geological Survey Paper 2015-1, 59-86.

 

Febbo, G.E., Friedman, R.M., Kennedy, L.A., and Nelson, J.L. (2019a). U-Pb geochronology of the Mitchell deposit, northwestern British Columbia. British Columbia Ministry of Energy, Mines and Petroleum Resources, British Columbia Geological Survey, GeoFile 2019-03, 8 p.

 

Febbo, G.E., Kennedy, L.A., Nelson, J.L., Savell, M.J., Campbell, M.E., Creaser, R.A., Friedman, R.M., van Straaten, B.I., and Stein, H.J. (2019b). The evolution and structural modification of the supergiant Mitchell Au-Cu porphyry, northwestern British Columbia. Economic Geology, 114(2), 303-324.

 

Fowler B., and Wells, R. (1995). The Sulphurets Gold zone, northwestern British Colombia. In Schroeter, T.G., ed., Porphyry deposits of the northwestern Cordillera of North America: Canadian Institute of Mining, Metallurgy and Petroleum, Special Volume 46, PART B - Porphyry Copper (±Au±Mo) Deposits of the Calc-Alkalic Suite, 484-492.

 

G&T (2007). Preliminary Assessment on Mitchell Zone Samples (KM 1909). June 2007.

 

G&T (2008). Pre-Feasibility Metallurgical Testing Mitchell Zone – Kerr Sulphurets (KM 2153). September 2008.

 

G&T (2009). Metallurgical and Pilot Plant Testing on Samples from the Kerr-Sulphurets-Mitchell (KSM) Project (KM 2344). December 2009.

 

G&T (2010). Ancillary Testing Kerr Sulphurets-Mitchell-KSM Project (KM 2755). December 2010.

 

G&T (2010). Bench Scale and Pilot Plant Testing Kerr-Sulphurets-Mitchell Project (KM 2670), August 2010.

 

G&T (2010). Miscellaneous Metallurgical Testing on Samples From the KSM Project (KM 2535). March 2010.

 

G&T (2011). Data Reports (KM 2897). February – March 2011.

 

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G&T (2011). Flotation and Cyanidation Testing – Mitchell Composites Years 0-10/0-20 – (KM 3081). November 28, 2011.

 

G&T (2011). Preliminary Metallurgical Testing on Samples from the Iron Cap Zone – KSM Project (KM 2748). January 2011.

 

G&T (2012). Flotation and CIL Cyanidation Testing – Kerr-Sulphurets-Mitchell Project – (KM 3080). January 10, 2012.

 

G&T (2012). Ore Hardness and Flotation Testing – Kerr-Sulphurets-Mitchell (KSM) Project – (KM 3174). January 6, 2012.

 

Golder (2011). Block Cave Mine Study. Report Number REP 0516_11. Submitted May17, 2011.

 

Golder (2012). Pre-Feasibility Block Cave Mine Design – Mitchell Deposit. Prepared for Seabridge Gold Inc. Submitted May 2012.

 

Gustafson, L.B., and Hunt, J.P. (1975). The porphyry copper deposit at El Salvador, Chile. Economic Geology, v.70, 857-912.

 

Hazen (2008). Comminution Testing (Project # 10724). February 2008.

 

IGRB (2016). Review of Water Dam, Water Management, and Tailings Storage Systems, KSM Project, British Columbia, Canada. Rev C.1. April 2016.

 

KCB (2009). 2008 KSM Site Investigation Report.

 

KCB (2010). 2009 KSM Site Investigation Report.

 

KCB (2011). 2010 KSM Site Investigation Report.

 

KCB (2012a), 2012 Engineering Design Update of Tailing Management Facility, December 21, 2012.

 

KCB (2012b). 2012 Site Investigation Report for the Mine Area.

 

KCB (2012c). 2012 TMF Site Investigations.

 

KCB (2012d). Engineering Design Update of Tailing Management Facility.

 

KCB (2012e). KSM Water Storage Dam - Value Engineering Study Report.

 

KCB (2013a). 2012 Geotechnical Design of Rock Storage Facilities and Design of Associated Water Management Facilities (Technical Report in Support of 2012 PFS), January 29, 2013.

 

KCB (2013b). KSM Project. Rock Storage Facilities Design Report. July 8, 2013.

 

KCB (2014). KSM Project Mitchell/McTagg RSF Water Collection for Selenium Treatment. March 4, 2014.

 

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KCB (2016a). Best Available Technology (BAT) Study for Tailing Management at the KSM Project. June 22, 2016.

 

KCB (2016b), 2016 Pre-Feasibility Study Update, Addendum Report, Tailing Management Facility Design – Rev. 1, September 22, 2016

 

KCB (2016c), 2016 Pre-Feasibility Study Update, Addendum Report – Mine Area Water Management – Rev. 1, September 27, 2016

 

KCB (2016d), 2016 Review of Regional Climate, 2016 Pre-Feasibility Study Update Addendum Report Tailing Management Facility Design – Rev. 1, September 22, 2016

 

KCB (2019a), September 2018 Site Visit TMF West Till Borrow Area Sampling and Geotechnical Laboratory Testing - Rev. 1, January 23, 2019.

 

KCB (2019b), Evaluation of Results from Batch 1 and 2 Monzonite Geochemical and Geotechnical Laboratory Testing Programs, March 20, 2019.

 

KCB (2020a) KSM 2020 Preliminary Economic Assessment, Review of PEA Mine Area Catchments – Rev. 1, February 12, 2020

 

KCB (2020b) KSM 2020 Preliminary Economic Assessment, Water and Waste Management - Basis of Quantities – Rev. 3, March 13, 2020

 

KCB (2020c) KSM 2020 Preliminary Economic Assessment, Diversion Tunnel Modifications, January 28, 2020

 

KCB (2021), “2021 Site Investigation Factual Report”. Prepared for KSM Mining ULC. February 2.

 

KCB (2022a), “2022 PFS Mine Area Water Tunnels Mitchell and McTagg Diversion Tunnel Design Final”. Prepared for KSM Mining ULC. June 7.

 

KCB (2022b), “2022 PFS Mine Area Water Tunnels Design”. Prepared for KSM Mining ULC. July 22.

 

Kirkham, R.V. (1963). The geology and mineral deposits in the vicinity of the Mitchell and Sulphurets glaciers, Northwest British Columbia. Unpublished M.Sc. thesis: Vancouver, Canada, University of British Columbia, 142 p.

 

Kirkham, R.V., and Margolis, J. (1995). Overview of the Sulphurets area, northwestern British Columbia. In Schroeter, T.G., ed., Porphyry deposits of the northwestern Cordillera of North America: Canadian Institute of Mining, Metallurgy and Petroleum, Special Volume 46, PART B - Porphyry Copper (±Au±Mo) Deposits of the Calc-Alkalic Suite, pp. 473-483.

 

Köeppern Machinery Australia Pty Ltd. (2010). High Pressure Comminution Test Work on Processing of Mitchell Zone Ore. February 2010.

 

MacDonald, A.J. (1993). The Iskut River Area, Northwestern British Columbia, Canada: an application of research in metallogenesis to enhance exploration success. In: MDRU Iskut River Metallogeny Project – Annual Report, Year 3, 7.1-7.37.

 

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Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

Margolis, J. (1993). Geology and intrusion-related copper-gold mineralization, Sulphurets, British Columbia. Unpublished Ph.D. thesis, Eugene, USA, University of Oregon, 289 p.

 

Metal Mine Effluent Regulations, SOR/2002-222.

 

Mines Act, RSBC. C. 293. (1996b).

 

National Instrument 43-101 Standards of Disclosure for Mineral Projects, June 30, 2011.

 

Nelson, J. and Colpron, M. (2007). Tectonics and metallogeny of the British Columbia, Yukon and Alaskan Cordillera, 1.8 Ga to the present. Mineral Deposits of Canada: A Synthesis of Major Deposit-Types, District Metallogeny, the Evolution of Geological Provinces, and Exploration Methods. Edited by W.D. Goodfellow. Geological Association of Canada, Mineral Deposits Division, Special Publication, 5, 755-791.

 

NMT (2022), MTT Automated Rail Transportation System. April 29, 2022.

 

Palacios, C., Hérail, G., Townley, B., Maksaev, V., Sepúlveda, F., de Parseval, P., Rivas, P., Lahsen, A. and Parada, M.A. (2001). The composition of gold in the Cerro Casale gold-rich porphyry deposit, Maricunga belt, northern Chile. The Canadian Mineralogist, 39(3), 907-915.

 

Piteau Associates (1991). Mine Rock and Overburden Piles Investigation and Design Manual: Interim Guidelines. Prepared for the British Columbia Mines Waste Rock Pile Research Committee by Piteau Associates Engineering Ltd. http://www.empr.gov.bc.ca/Mining/Permitting-Reclamation/Geotech/Documents/MinedRock+OverburdenPiles/MinedRockOverburdenPile_Investigation+DesignManual.pdf (accessed January 2013).

 

Placer Dome Inc. (1990). Metallurgical Research Centre, Kerr Project Report No.1. October 1990.

 

Placer Dome Inc. (1991). Metallurgical Research Centre, Kerr Project Report No.2. April 1991.

 

Pocock (2009). Flocculant Screening, Gravity Sedimentation, Pulp Rheology, Vacuum Filtration and Pressure Filtration Studies. December 2009.

 

Rescan (2013). Application for an Environmental Assessment Certificate/Environmental Impact Statement for the KSM Project. May 2013.

 

SGS Minerals Services (2010). An Investigation into the Grindability and Flotation Characteristics of Two Samples From the KSM Deposit. February 2010.

 

SGS Minerals Services (2010). An Investigation into the Recovery of Cyanide and Detoxification of Leach Tailing From Cyanidation of KSM Project Samples. February 2010.

 

Sillitoe, R.H. (2010). Porphyry copper systems. Economic geology, 105(1), 3-41.

 

Singer, D.A., Berger, V.I., and Moring, B.C. 2008. Porphyry copper deposits of the world: Database and grade and tonnage models. U.S. Geological Survey Open-File Report 2008-1155.

 

SRK (2022), “KSM_MTT TER_FINAL DRAFT Report_CAPR000860_20220421.pdf”

 

Seabridge Gold Inc.27-6219221-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

Surface Science Western, the University Of Western Ontario (2010). Sub-microscopic Gold determination of Cyanide Leach Feeds. December 21, 2010.

 

Surface Science Western, the University Of Western Ontario (2016). Sub-microscopic Gold Determination of Leach Residues. August 30, 2016.

 

Tetra Tech (2012). 2012 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study. Document No. 1252880100-REP-R0001-05. Prepared for Seabridge Gold Inc. June 22, 2012.

 

Tetra Tech (2014). Feasibility Study and Technical Report Update on the Brucejack Project, Stewart, BC. Document No. 1491990100-REP-R0001-01. June 19, 2014.

 

Tetra Tech (2016). 2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment. Document No. 735-1552880100-REP-R0002-02. Prepared for Seabridge Gold Inc. October 6, 2016.Wardrop (2008). Kerr-Sulphurets-Mitchell Preliminary Economic Assessment 2008, report written for Seabridge and later filed as NI 43-101 Technical Report, December 19, 2008.

 

Tetra Tech (2020). KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update. Document No. 219221-01-RPT-002. Prepared for Seabridge Gold Inc. April 3, 2020.

 

W.N. Brazier and Associates Inc., KSM Mining ULC 2019 Report on Fuel Prices ,Rev. 1. October 17 , 2019.

 

W.N. Brazier and Associates Inc., KSM Mining ULC KSM 2019 PFS Update Power Supply – NTL Distribution, Rev. 2. October 17, 2019.

 

W.N. Brazier and Associates Inc., KSM Mining ULC KSM Project Cost of Electric Power 170,000 TPD 2019 PEA, Rev. 0. January 12, 2020.

 

Wardrop (2011). Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study Update 2011, report written for Seabridge and later filed as NI 43-101 Technical Report, June 15, 2011.

 

Water Act, RSBC. C. 483. (2003).

 

WSP Golder (2022), “2021 Hydrogeological Site Investigations”. Prepared for KSM Mining ULC. February 8.

 

Seabridge Gold Inc.27-7219221-01-RPT-001

KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and
Preliminary Economic Assessment

NI 43-101 Technical Report

  

 

 

 

Certificates of Qualified Persons

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

 

Certificate of Qualified Person

 

 

I, Hassan Ghaffari, P.Eng., do hereby certify:

 

I am a Director of Metallurgy with Tetra Tech Inc. with a business address at Suite 1000, 10th Floor, 885 Dunsmuir Street, Vancouver, BC, V6C 1N5.

 

This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report”, with an effective date of August 08, 2022 (the “Technical Report”).

 

I am a graduate of the University of Tehran (M.A.Sc., Mining Engineering, 1990) and the University of British Columbia (M.A.Sc., Mineral Process Engineering, 2004).

 

I am a member in good standing of the Engineers and Geoscientists British Columbia (#30408).

 

My relevant experience includes 31 years of experience in mining and mineral processing plant operation, engineering, project studies and management of various types of mineral processing, including hydrometallurgical mineral processing for porphyry mineral deposits.

 

I am a “Qualified Person” for the purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

 

I conducted a personal inspection of the KSM property on September 20, 2014.

 

I am responsible for Sections 1.1, 1.3, 1.11, 1.13, 1.14.1, 1.14.7 (except mining and TMF), 1.14.8, 1.15, 1.16, 2.0, 3.0, 5.0, 18.1, 18.5, 18.6, 18.7, 18.8, 18.9, 18.12, 18.13, 18.14, 18.15, 21.1 and 21.2 (except mining, tunnels, rail system, permanent electrical power supply and distribution, energy recovery and mini hydro generation station costs), 22.0, 24.18 (except mining, tunnels, rail system, TMF, water mgmt., permanent electrical power supply and distribution and energy recovery), 24.21 (except mining, tunnels, rail system, permanent electrical power supply and distribution, energy recovery and mini hydro generation station costs), 24.22, 25.1, 25.2.6, 25.3.4, 26.1 and 27.0 (only references from sections for which I am responsible) of the Technical Report.

 

I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

 

I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study” dated June 22, 2012, “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016, and “2020 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update” dated April 30, 2020.

  

I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

 

As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

 

Signed and dated this 8th day of August, 2022

 

  “signed and stamped”  

Hassan Ghaffari, P.Eng.
Director of Metallurgy
Tetra Tech Inc.  

 

 

 

 

Certificate of Qualified Person

 

 

I, Jianhui (John) Huang, Ph.D., P.Eng., do hereby certify:

 

I am a Senior Metallurgist with Tetra Tech Inc. with a business address at Suite 1000, 10th Floor, 885 Dunsmuir Street, Vancouver, British Columbia, V6C 1N5.

 

This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report”, with an effective date of August 08, 2022 (the “Technical Report”).

 

I am a graduate of North-East University, China (B.Eng., 1982), Beijing General Research Institute for Non-ferrous Metals, China (M.Eng., 1988), and Birmingham University, United Kingdom (Ph.D., 2000).

 

I am a member in good standing of the Engineers and Geoscientists British Columbia (#30898).

 

My relevant experience includes over 36 years involvement in mineral processing for base metal ores, gold and silver ores, and rare metal ores, and mineral processing plant operation and engineering including hydrometallurgical mineral processing for porphyry mineral deposits.

 

I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

 

My most recent personal inspection of the KSM property was from June 21, 2017 to June 22, 2017. I visited ALS Metallurgical laboratory several times during the last 10 years for the KSM project, including the most visit on November 02, 2021, to witness metallurgical testing and samples.

 

I am responsible for Sections 1.8, 1.12, 1.14.6, 1.14.7 (excluding mining costs), 13.0, 17.0, 19.0, 21.2.3, 21.2.4, 21.3 (excluding mining costs), 24.17, 24.19, 24.21.2 (excluding open pit mining and underground mining costs) 25.2.4, 25.3.2, 25.4 and 25.5 (metallurgy and mineral process only), 26.2.3 and 27.0 (only references from sections for which I am responsible) of the Technical Report.

 

I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

 

I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment 2008”, dated December 22, 2008, the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment Addendum – 2009” dated September 8, 2009, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study” dated March 31, 2010, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study Update” dated June 15, 2011, “2012 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study” dated June 22, 2012, “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016, and “2020 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update” dated April 30, 2020.

  

I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

 

As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

 

Signed and dated this 8th day of August, 2022

 

  “signed and stamped”  

Jianhui (John) Huang, Ph.D., P.Eng.
Senior Metallurgist
Tetra Tech Inc.  

 

 

 

  

Certificate of Qualified Person

 

 

I, Henry H. Kim, P.Geo., do hereby certify:

 

I am employed as a Senior Resource Geologist with Wood Canada Limited, with a business address at #400-111 Dunsmuir St., Vancouver, BC, Canada V6B 5W3.

 

This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report”, with an effective date of August 08, 2022 (the “Technical Report”).

 

I graduated with a B.Sc. degree in geology from University of British Columbia in 2008. I completed the Applied Geostatistics Citation Program, with the University of Alberta in 2014.

 

I am a registered Professional Geoscientist with the Engineers and Geoscientists British Columbia (#42519).

 

I have practiced my profession for 14 years. I have been involved in exploration drilling programs involving core logging, sampling, QAQC, and database validation. I conducted onsite grade control and management of mine operation crews for an open pit mine in eastern Canada. I have conducted audits and due diligence exercises on geological models, sampling databases, drill hole spacing studies, preparation of resource models, validation of mineral resource estimates, and mineral resource estimates on advanced mining studies and active mine operations.

 

I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

 

I conducted a personal inspection of the KSM property from May 24 to May 26 in 2022.

 

I am responsible for Sections 1.2, 1.4, 1.5, 1.6, 3.0 (matters related to mineral claims and royalties), 4.0, 6.0, 7.0, 8.0, 9.0, 10.0, 11.0, 12.0, 14.0, 23.0, 25.2.1, 25.5 (matters related to mineral resources), 25.6, and 26.0 (matters related to mineral resources) and 27.0 (only references from sections for which I am responsible) of the Technical Report.

 

I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

 

I have no prior involvement with the KSM Property that is the subject of the Technical Report.

 

I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

  

As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

 

Signed and dated this 8th day of August, 2022

 

  “signed and stamped”  

Henry H. Kim, P.Geo.
Senior Resource Geologist
Wood Canada Limited

 

 

 

 

Certificate of Qualified Person

 

 

I, James H. Gray, P.Eng., do hereby certify:

 

I am a Mining Engineer with Moose Mountain Technical Services with a business address at #210 1510 2nd Street North, Cranbrook, British Columbia, V1C 3L2.

 

This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report”, with an effective date of August 08, 2022 (the “Technical Report”).

 

I am a graduate of the University of British Columbia, (Bachelor of Applied Science – Mineral Engineering, 1975).

 

I am a member in good standing of Engineers and Geoscientists British Columbia (#11919).

 

My relevant experience includes mine operation, supervision, and mining engineering in North America, South America, Australia, Eastern Europe, and Greenland.

 

I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

 

My most recent personal inspection of the KSM property was from June 21 to 22, 2017.

 

I am responsible for MTT and rail systems for 1.7, 1.9.4 (MTT), 1.9.5, 1.11 and 1.12 (open pit mining, tunnels and rail system), 1.14.2, 1.14.3, 1.14.5, 1.14.7 (open pit mining, tunnels and rail system), 15.0, 16.0, 18.3, 18.4, 21.0 (open pit mining, tunnels and rail system), 24.0 (open pit mining, MTT, tunnel costs and rail system), 25.3.1, 25.0 and 26.0 (open pit mining, MTT and rail system) and 27.0 (only references from sections for which I am responsible) of the Technical Report.

 

I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

 

I have had previous involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment 2008”, dated December 22, 2008, the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment Addendum – 2009” dated September 8, 2009, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study” dated March 31, 2010, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study Update” dated June 15, 2011, the “2012 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study” dated June 22, 2012 and amended November 11, 2014, and the “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016, and “2020 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update” dated April 30, 2020.

 

I have read the NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with the NI 43-101.

 

As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

 

Signed and dated this 8th day of August, 2022

 

  “signed and stamped”  

James H. Gray, P.Eng.
Principal Mining Engineer
Moose Mountain Technical Services

 

 

 

 

  

Signed and dated this 8th day of August, 2022

 

  “signed and stamped”  

Ross D. Hammett, Ph.D., P.Eng.
Senior Engineer & Principal
WSP Golder Inc.

 

 

 

  

Certificate of Qualified Person

 

 

I, Derek Kinakin, P.Geo., P.G., do hereby certify:

 

I am a Senior Engineering Geologist with BGC Engineering Inc. with a business address at 234 St. Paul Street, Kamloops, British Columbia, V2C 6G4.

 

This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report”, with an effective date of August 08, 2022 (the “Technical Report”).

 

I am a graduate of Simon Fraser University (B.Sc. (Hons), 2002; M.Sc., 2005).

 

I am a member in good standing of Engineers and Geoscientists British Columbia (#32720).

 

My relevant experience includes 20 years of rock mechanics research, slope stability assessments, and slope designs for open pit mines in Canada, US, and Africa.

 

I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

 

My most recent personal inspection of the KSM property was August 13, 2018 to August 17, 2018.

 

I am responsible for Section 1.9.1 (geohazards), 16.2.4 (pit slope geotech.), 25.4 (geohazards), 26.2.1 (pit slope geotech.) and 27.0 (only references from sections for which I am responsible) of the Technical Report.

 

I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

 

I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016, and “2020 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update” dated April 30, 2020.

 

I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

 

As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

  

Signed and dated this 8th day of August, 2022

 

  “signed and stamped”  

Derek Kinakin, P.Geo., P.G.
Senior Engineering Geologist
BGC Engineering Inc.

 

 

 

 

Certificate of Qualified Person

 

 

I, David A. Willms, P.Eng., do hereby certify:

 

I am a Senior Geotechnical Engineer with Klohn Crippen Berger Ltd. with a business address at Suite 500, 2955 Virtual Way, Vancouver, British Columbia, Canada, V5M 4X6.

 

This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report”, with an effective date of August 08, 2022 (the “Technical Report”).

 

I am a graduate of the University of British Columbia, 2003, with a Bachelor of Applied Science in Geological Engineering, and 2010 with an M.Eng. in Geological Engineering.

 

I am a member in good standing of the Engineers and Geoscientists British Columbia (#33062).

 

My relevant experience includes more than 18 years of experience in the mining industry, with a focus on tailings management.

 

I am a “Qualified Person” for purposes of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

 

I conducted a personal inspection of the KSM property on May 24, 2022.

 

I am responsible for Section 1 (RSF, WSF, TMF, site water management and water tunnels), 18.2, 21.0 and 24.0 (RSF, WSF, TMF, site water management and water tunnels), 25.0 and 26.0 (RSF, WSF, TMF, site water management, water tunnels and site geotechnical) and 27.0 (only references from sections for which I am responsible) of the Technical Report.

 

I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

 

I have no prior involvement with the KSM Property that is the subject of the Technical Report.

 

I have read NI 43-101 and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

 

As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

  

Signed and dated this 8th day of August, 2022

 

  “signed and stamped”  

David A. Willms, P. Eng.
Senior Geotechnical Engineer
Klohn Crippen Berger Ltd.

 

 

 

 

Certificate of Qualified Person

 

 

I, Neil Brazier, P.Eng., do hereby certify:

 

I am a Principal with WN Brazier Associates Inc. with a business address at #8–3471 Regina Ave., Richmond, BC. V6X 2K8

 

This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report”, with an effective date of August 08, 2022 (the “Technical Report”).

 

I am a graduate of the University of Saskatchewan (B.Sc. Electrical Engineering, 1969).

 

I am a member in good standing of the Engineers and Geoscientists British Columbia (#8337).

 

My relevant experience includes engineering, construction supervision, and commissioning of a large number of diesel and combustion turbine power plants, high-voltage transmission lines and substations for mining applications.

 

I am a “Qualified Person” for purposes of National Instrument 43-101 (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

 

My most recent personal inspection of the KSM property was from October 4 to October 7 in 2021.

 

I am responsible for portions of Sections 1.11 and 1.14.7 related to permanent electrical power supply and distribution, energy recovery, and mini hydro generation station costs of the Technical Report, 18.10, 18.11, portions of Section 21.1 and 21.2 related to permanent electrical power supply and distribution, energy recovery, and mini hydro generation station costs of the Technical Report, portions of Section 24.18.2, 24.18.3, 24.18.4 and portions of Section 24.21 related to permanent electrical power supply and distribution and 27.0 (only references from sections for which I am responsible) of the Technical Report.

 

I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

 

I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment 2008”, dated December 22, 2008, the “Kerr-Sulphurets-Mitchell Preliminary Economic Assessment Addendum – 2009” dated September 8, 2009, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study” dated March 31, 2010, the “Kerr-Sulphurets-Mitchell (KSM) Prefeasibility Study Update” dated June 15, 2011, the “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study” dated June 22, 2012 and the “2016 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update and Preliminary Economic Assessment” dated October 6, 2016, and “2020 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update” dated April 30, 2020.

 

I have read NI 43-101 Standards of Disclosure for Mineral Projects (NI 43–101) and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

 

As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

 

Signed and dated this 8th day of August, 2022

 

  “signed and stamped”  

Neil Brazier, P.Eng.
Principal
WN Brazier Associates Inc.

 

 

 

  

Certificate of Qualified Person

 

 

I, Rolf Schmitt, P.Geo., do hereby certify:

 

I am a Technical Director with ERM Consultants Canada Ltd. with a business address at 1500-1111 West Hastings St., Vancouver, British Columbia, V6E 2J3.

 

This certificate applies to the technical report entitled “KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study and Preliminary Economic Assessment, NI 43-101 Technical Report”, with an effective date of August 08, 2022 (the “Technical Report”).

 

I graduated with a Master of Science in Geology from the University of Ottawa in 1993. In addition, I have obtained a Master of Science, Regional Planning from the University of British Columbia, 1985, and Honours Bachelor of Science, Geology, from the University of British Columbia in 1977.

 

I am a member in good standing of the Engineers and Geoscientists British Columbia (#19824).

 

I have worked as a geologist, project manager, technical director and senior policy advisor for a total of 44 years since my graduation from university. Key areas of experience of relevance to the KSM project include: Senior Exploration Geologist, Kidd Creek Mines (northwest BC), Research Exploration Geochemist (Geological Survey of Canada), Senior Land Use Geologist and Mineral Policy Specialist (BC Ministry of Energy, Mines and Petroleum Resources), Technical Director and Project Manager (Rescan and ERM), delivering mine Environmental Assessment Projects in BC, permitting major mines in BC and undertaking mine ESG Due Diligence assignments across Canada and internationally as NI 43-101 QP. I have worked on multiple mine assignments through all phases of life-of-mine as part of owners’ exploration/ engineering/ environmental teams and as a senior strategic advisor.

 

I am a “Qualified Person” for purposes of National Instrument 43-101 (NI 43-101) for those sections of the Technical Report that I am responsible for preparing.

 

My most recent personal inspection of the KSM property was on July 12, 2019.

 

I am responsible for Sections 1.10, 20.0, 24.20, 26.2.4 (water balance, temporary water treatment plants, and geochemistry database) and 27.0 (only references from sections for which I am responsible) of the Technical Report.

  

I am independent of Seabridge Gold Inc. as Independence is defined by Section 1.5 of NI 43-101.

 

I have had involvement with the KSM property that is the subject of the Technical Report, in acting as a Qualified Person for the “2020 KSM (Kerr-Sulphurets-Mitchell) Prefeasibility Study Update” dated April 30, 2020.

 

I have read NI 43-101 Standards of Disclosure for Mineral Projects (NI 43–101) and the sections of the Technical Report that I am responsible for have been prepared in compliance with NI 43-101.

 

As of the date of this certificate, to the best of my knowledge, information and belief, the section of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

 

Signed and dated this 8th day of August, 2022

 

  “signed and stamped”  

Rolf Schmidt, P.Geo.
Division Managing Director, Canada
ERM Consultants Canada Ltd.