EX-99.1

Prepared by:

In Collaboration with:

EX-99.1 2 d570829dex991.htm EX-99.1 EX-99.1

NI 43-101

Feasibility Study of the Rainy River Gold Project,

Ontario, Canada

LOGO

Prepared for:

NEW GOLD INC.

Suite 1800, Two Bentall Centre

555 Burrard Street,

Vancouver, BC

LOGO

Report Date (readdressed): July 31, 2013

Effective Date: April 10, 2013

Colin Hardie, P. Eng., BBA Inc.

David Runnels, Eng., BBA Inc.

Patrice Live, Eng., BBA Inc.

Sheila E. Daniel, M.Sc., P. Geo, AMEC

David G. Ritchie, P. Eng., AMEC

Adam Coulson, PhD., P. Eng., AMEC

Glen Cole, P.Geo. SRK Consulting (Canada) Inc.

Dorota El-Rassi, P. Eng., SRK Consulting (Canada) Inc.

Donald Tolfree, P. Eng., Golder Associates Ltd.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

DATE AND SIGNATURE PAGE

This Report is effective as of the 10^th day of April 2013 and readdressed to New Gold Inc. on July 31, 2013.


Original signed and sealed				July 31, 2013
Colin Hardie, P.Eng.				Date
Department Manager, Mining and Metals
BBA Inc.

Original signed and sealed				July 31, 2013
David Runnels, Eng.				Date
Project Manager, Mining and Metals
BBA Inc.

Original signed and sealed				July 31, 2013
Patrice Live, Eng.				Date
Manager of Mining
BBA Inc.

Original signed and sealed				July 31, 2013
Sheila E. Daniel, M.Sc., P.Geo.				Date
Head Environmental Management
Senior Associate Geoscientist
AMEC – Environment & Infrastructure

Original signed and sealed				July 31, 2013
David G. Ritchie, P.Eng.				Date
Geotechnical Engineering Group Manager
Senior Associate Geotechnical Engineer
AMEC – Environment & Infrastructure

Original signed and sealed				July 31, 2013
Adam Coulson, PhD., P.Eng.				Date
Rock Engineering Group Manager
Senior Associate Rock Mechanics Engineer
AMEC – Environment & Infrastructure

Original signed and sealed				July 31, 2013
Glen Cole, P.Geo.				Date
Principal Consultant (Resource Geology)
SRK Consulting (Canada) Inc.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project


Original signed and sealed				July 31, 2013
Dorota El-Rassi, P.Eng.				Date
Senior Consultant (Resource Geology)
SRK Consulting (Canada) Inc.

Original signed and sealed				July 31, 2013
Donald Tolfree, P.Eng.				Date
Mining Engineer
Golder Associates Ltd.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

CERTIFICATE

To accompany the technical report entitled: “Feasibility Study of the Rainy River Gold Project, Ontario, Canada” originally dated May 23, 2013 and effective April 10, 2013 and subsequently readdressed to New Gold Inc. on July 31, 2013 (the “Technical Report”).

I, Colin Hardie, P. Eng., as a co-author of the Technical Report, do hereby certify that:

1)	I am currently employed as Department Manager – Mining and Metals in the consulting firm BBA Inc.:

630 René-Lévesque Boulevard West

Suite 1900

Montreal, Quebec H3B 4V5 Canada

I graduated from the University of Toronto in 1996 with a BASc in Geological and Mineral Engineering. In 1999, I graduated from McGill University of Montreal with an M.Eng in Metallurgical Engineering and in 2008 obtained a Master of Business Administration (MBA) degree from the University of Montreal (HEC);

I am a member in good standing of the Professional Engineers of Ontario (Member Number: 90512500) and of the Canadian Institute of Mining, Metallurgy, and Petroleum (Member Number: 140556). I have practiced my profession continuously since my graduation. I have been employed in mining operations, consulting engineering and applied metallurgical research for over 15 years;

4)	I visited the site on June 16, 2011;

I have read the definition of “qualified person” set out in the National Instrument 43-101 - Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that, by reason of my education, affiliation with a professional association, and past relevant work experience, I fulfill the requirements to be an independent qualified person for the purposes of NI 43-101;

6)	I am independent of the issuer, New Gold Inc., as defined in Section 1.5 of NI 43-101;

7)	I am responsible for sections 1, 2, 3, 4, 19, 21 (except 21.4, 21.5, 21.15.2 and 21.15.3), 22, 23, 25, 26, 27, Appendix A and Appendix B of the Technical Report;

I have had prior involvement with the subject property having co-authored a previous technical report entitled “Preliminary Economic Assessment of the Rainy River Gold Property” prepared by BBA in December 2011; and the “Preliminary Economic Assessment Update of the Rainy River Gold Property” prepared by BBA in October 2012;

9)	I have read NI 43-101 and the sections of the Technical Report under my responsibility have been prepared in compliance therewith; and

10)	That, as of the effective date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report under my responsibility contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

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NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

CERTIFICATE

I, David Runnels, Eng., as a co-author of the Technical Report, do hereby certify that:

1)	I am the Project Manager – Mining and Metals with the firm BBA Inc. with an office at 630 René- Lévesque Blvd. West, Suite 1900, Montréal, Quebec, H3B 4V5 Canada;

2)	I am a graduate of the Queen’s University, Kingston Ontario, Canada with a B. Sc. In Metallurgy in 1971. I have practiced my profession continuously since my graduation from university;

3)	I am a registered member in good standing of the Order of Engineers of Québec (#22450) and I am a member of the Canadian Institute of Mining;

4)	I visited the site on October 11, 2012;

I have read the definition of “qualified person” set out in the NI 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that, by reason of my education, affiliation with a professional association, and past relevant work experience, I fulfill the requirements to be an independent qualified person for the purposes of NI 43-101;

6)	I am independent of the issuer, New Gold Inc., as defined in Section 1.5 of NI 43-101;

7)	I am responsible for Sections 13, 17, 18 (except 18.2 and 18.10), 24 and Appendix H of the Technical Report;

I have had prior involvement with the subject property having contributed to a previous technical report entitled “Preliminary Economic Assessment of the Rainy River Gold Property” prepared by BBA in December 2011; and the “Preliminary Economic Assessment Update of the Rainy River Gold Property” prepared by BBA in October 2012;

9)	I have read NI 43-101 and the sections of the Technical Report I am responsible for and confirm that the sections under my responsibility in the Technical Report has been prepared in compliance therewith; and

10)	That, as of the effective date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report under my responsibility contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

CERTIFICATE

To accompany the technical report entitled: “Feasibility Study of the Rainy River Gold Project, Ontario Canada” originally dated May 23, 2013 and effective April 10, 2013 and subsequently readdressed to New Gold Inc. on July 31, 2013 (the “Technical Report”).

I, Patrice Live, Eng., as a co-author of the Technical Report, do hereby certify that:

1)	I am currently employed as Manager of Mining in the consulting firm BBA Inc.: 630 René-Lévesque Boulevard West, Suite 1900, Montreal, Quebec H3B 4V5 Canada;

2)	I graduated from the University of Laval (Québec City) in 1976 with a BASc in Mining Engineering. I have worked as a mining engineer continuously since my graduation from university;

I am a member in good standing of the Order of Engineers of Québec (#38991) and I am a member of the Canadian Institute of Mining. Experience includes a full range of studies from: preliminary economic assessments, prefeasibility studies and feasibility studies, as well as due diligence audits and technical reviews, ore reserves estimation, mining methods, mine design and development, production scheduling and planning, equipment sizing, capital and operating cost estimates with cash flow models, and financial analysis;

4)	I visited the site on October 11, 2012;

I have read the definition of “qualified person” set out in the National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that, by reason of my education, affiliation with a professional association, and past relevant work experience, I fulfill the requirements to be an independent qualified person for the purposes of NI 43-101;

6)	I am independent of the issuer, New Gold Inc., as defined in Section 1.5 of NI 43-101;

7)	I am responsible for Sections 15 (except 15.2), 16 (except 16.2.1 and 16.3), 21.4 and 21.15.2 of the Technical Report;

9)	I have read NI 43-101 and the sections of the Technical Report under my responsibility have been prepared in compliance therewith; and

10)	That, as of the effective date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report under my responsibility contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

CERTIFICATE

I, Sheila Ellen Daniel, M.Sc., P.Geo. do hereby certify that:

1)	I am Head Environmental Management, Senior Associate Geoscientist, in the consulting firm:

AMEC Environment & Infrastructure, a Division of AMEC Americas Limited

160 Traders Blvd. East, Suite 110

Mississauga, ON

Canada L4Z 3K7;

2)	I graduated from McMaster University in 1990 with a M.Sc. and University of Western Ontario with a B.Sc. (Honours); I have practiced my profession for twenty-two years since my graduation from university;

3)	I am Professional Geoscientist in the Province of Ontario (Reg. # 0151);

4)	I visited the site on May 19, 2011;

6)	I am independent of the issuer, New Gold Inc., as defined in Section 1.5 of NI 43-101;

7)	I am responsible for Section 20 of the Technical Report;

I have had prior involvement with the subject property having contributed input on environmental aspects to a previous technical report entitled “Preliminary Economic Assessment of the Rainy River Gold Property” prepared by BBA in December 2011; and the “Preliminary Economic Assessment Update of the Rainy River Gold Property” prepared by BBA in October 2012;

9)	I have read NI 43-101 and confirm that the sections of the Technical Report under my responsibility have been prepared in compliance therewith; and

10)	That, as of the effective date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report under my responsibility contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

CERTIFICATE

I, David G. Ritchie, P.Eng. do hereby certify that:

1)	I am a Senior Associate Geotechnical Engineer and Geotechnical Engineering Group Manager in the consulting firm:

AMEC Environment & Infrastructure, a Division of AMEC Americas Limited

160 Traders Blvd. East, Suite 110

Mississauga, ON L4Z 3K7

Canada

2)	I graduated from Ryerson Polytechnic University in 1995 with a B.Eng. in Civil Engineering and in 2000 from the University of Western Ontario with a M.Eng.

3)	I am Professional Engineer in the Province of Ontario (Reg. # 90488198). I have practiced my profession for eighteen years since my graduation from university;

4)	I visited the site on June 22, 2011;

6)	I am independent of the issuer, New Gold Inc., as defined in Section 1.5 of NI 43-101;

7)	I am responsible for Sections 18.2, 18.10 and 16.2.1.2 of the Technical Report;

I have had prior involvement with the subject property having contributed to a previous technical report entitled: “Preliminary Economic Assessment of the Rainy River Gold Property” prepared by BBA in December 2011; and the “Preliminary Economic Assessment Update of the Rainy River Gold Property”, prepared by BBA in October 2012;

9)	I have read NI 43-101 and the parts of the Technical Report under my responsibility have been prepared in compliance therewith; and

10)	That, as of the date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report under my responsibility contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

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NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

CERTIFICATE

I, Adam Coulson, Ph.D., P. Eng., as a co-author of the Technical Report, do hereby certify that:

1)	I am currently employed as a Senior Associate Rock Mechanics Engineer and Rock Engineering Group Manager in the consulting firm AMEC Environment & Infrastructure, a division of AMEC Americas Ltd.:

160 Traders Blvd. E.,

Suite 110

Mississauga, Ontario L4Z 3K7 Canada

2)	I graduated with a B.Eng, from Camborne School of Mines, UK in 1990; obtained a MSc. (Eng) from Queens University, Canada in 1996; and a Ph.D. from the University of Toronto, Canada in 2009;

I am a member in good standing of the Professional Engineers of Ontario (Member No. 100049242) and of the Canadian Institute of Mining, Metallurgy, and Petroleum (Member No. 146473). I have practiced my profession continuously since my graduation. I have been employed in mining operations, consulting engineering and rock mechanics research for over 22 years;

4)	I visited the site on January 24 to 27, 2012;

6)	I am independent of the issuer, New Gold Inc., as defined in Section 1.5 of NI 43-101;

7)	I am responsible for sections 16.2.1.1 and 16.3.1 of the Technical Report;

8)	I have had no prior involvement with the subject property prior to the commencement of the work required for the Technical Report;

9)	I have read NI 43-101 and the sections of the Technical Report under my responsibility have been prepared in compliance therewith; and

10)	That, as of the effective date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report under my responsibility contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

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NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

CERTIFICATE

I, Glen Cole residing at 15 Langmaid Court, Whitby, Ontario do hereby certify that:

1)	I am a Principal Consultant (Resource Geology) with the firm of SRK Consulting (Canada) Inc. (SRK) with an office at Suite 1300, 151 Yonge Street, Toronto, Ontario, Canada;

I am a graduate of the University of Cape Town in South Africa with a B.Sc (Hons) in Geology in 1983; I obtained an M.Sc (Geology) from the University of Johannesburg in South Africa in 1995 and an M.Eng in Mineral Economics from the University of the Witwatersrand in South Africa in 1999. I have practiced my profession continuously since 1986. Since 2006, I have estimated and audited mineral resources for a variety of early and advanced base and precious metals projects in Africa, Canada, Chile and Mexico. Between 1989 and 2005 I have worked for Goldfields Ltd at several underground and open pit mining operations in Africa and held positions of Mineral Resources Manager, Chief Mine Geologist and Chief Evaluation Geologist, with the responsibility for estimation of mineral resources and mineral reserves for development projects and operating mines;

3)	I am a Professional Geoscientist registered with the Association of Professional Geoscientists of the Province of Ontario (APGO#1416) and am also registered as a Professional Natural Scientist with the South African Council for Scientific Professions (Reg#400070/02);

4)	I have personally inspected the Rainy River Project and surrounding areas on a few occasions, most recently from March 24 to 26, 2010;

I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional association and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101;

6)	I am independent of the issuer, New Gold Inc., as defined in Section 1.5 of NI 43-101;

7)	I am responsible for sections 4, 5, 6, 7, 8, 9, 10, 11, 12 and Appendices C to G of the Technical Report;

I have had prior involvement with the subject property having co-authored previous technical reports prepared by SRK in April 2009, April 2011 and April 9, 2012 (amended June 4, 2012) and a mineral resource model in February 2010. I also contributed to a technical report entitled: “Preliminary Economic Assessment of the Rainy River Gold Property” prepared by BBA in December 2011; and the “Preliminary Economic Assessment Update of the Rainy River Gold Property” prepared by BBA in October 2012;

9)	I have read NI 43-101 and the sections of the Technical Report under my responsibility have been prepared in compliance therewith; and

10)

That, as of the effective date of the Technical Report, to the best of my knowledge, information and belief, the sections of the Technical Report for which I am responsible for contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

CERTIFICATE

I, Dorota El-Rassi, residing at 70 Portsdown Road, Scarborough, Ontario do hereby certify that:

1)	I am a Senior Consultant (Resource Geology) with the firm of SRK Consulting (Canada) Inc. with an office at Suite 1300, 151 Yonge Street, Toronto, Ontario, Canada;

I am a graduate of the University of Toronto with a BA.Sc (Hons) in 1997 and a MSc. in Geology in 2000. I have practiced my profession continuously since 1997. I have over 10 years’ experience in mineral exploration, resource estimation and consulting. Prior to joining SRK, I worked for Watts, Griffis and McOuat as a resource geologist. As a Resource Engineer, I estimated and audited projects in North America, South America, Asia and Africa. My experience includes gold, silver, copper, nickel, zinc, PGE and industrial mineral deposits. Areas of expertise are resource estimation, geological modelling and exploration project management;

3)	I am a Professional Engineer registered with the Association of Professional Engineers of the province of Ontario (Licence: 100012348) and a fellow with the Geological Association of Canada;

4)	I have not personally visited the project area but relied on a site visit completed by Mr. Glen Cole, P.Geo, a co-author of the Technical Report;

I have read the definition of “qualified person” set out in National Instrument 43-101- Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional association and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101;

6)	I am independent of the issuer, New Gold Inc., as defined in Section 1.5 of NI 43-101;

7)	I contributed towards Section 14 of the Technical Report and I accept professional responsibility that section of the Technical Report;

I have had prior involvement with the subject property having co-authored a previous technical reports prepared by SRK in April 2009, April 2011 and April 9, 2012 (amended June 4, 2012) and a mineral resource model in February 2010. I contributed to a previous technical report entitled “Preliminary Economic Assessment of the Rainy River Gold Property” prepared by BBA in December 2011; and the “Preliminary Economic Assessment Update of the Rainy River Gold Property” prepared by BBA in October 2012;

9)	I have read NI 43-101 and the section of the Technical Report I am responsible for and confirm that the Technical Report has been prepared in compliance therewith; and

10)	That, as of the effective date of the Technical Report, to the best of my knowledge, information and belief, the section of the Technical Report under my responsibility contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

CERTIFICATE

I, Donald Tolfree, of New Westminster, British Columbia, do hereby certify that:

•

I am a Mining Engineer with Golder Associates Ltd. with a business address at 500-4260 Still Creek Drive, Burnaby, B.C., V5C 6C6.

•

I am a graduate of McGill University, Mining Engineering (B. Eng. 2000) and University of British Columbia (MASc 2004). I have practiced my profession continuously since graduation. My relevant experience with respect to this deposit type includes four years working at a number of underground mining operations with major mining companies including Noranda (Brunswick Mine), Hemlo Gold (Golden Giant Mine) and Barrick (Bousquet 2) and three years participating in the completion of various pre-feasibility and feasibility studies while working at Golder Associates.

•

I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (License #32557).

•

I visited the site in October of 2012.

•

I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that, by reason of my education, affiliation with a professional association and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purpose of NI 43-101.

•

I am independent of the issuer, New Gold Inc., as defined in Section 1.5 of NI 43-101;

•

I am responsible for Sections 15.2, 16.3 (except 16.3.1), 21.5 and 21.15.3 of this Technical Report.

•

I have read NI 43-101 and the sections of the Technical Report under my responsibility have been prepared in compliance therewith; and

•

That, as of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report under my responsibility contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading;

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

CAUTIONARY NOTE WITH RESPECT TO FORWARD LOOKING INFORMATION

This document contains “forward-looking information” as defined in applicable securities laws (referred to herein as “forward-looking statements”). Forward looking statements include, but are not limited to, statements with respect to the cost and timing of the development of the Rainy River Gold project, the other economic parameters of the project, as set out in its feasibility study; the success and continuation of exploration activities; estimates of mineral reserves and resources; the future price of gold; government regulations and permitting timelines; estimates of reclamation obligations that may be assumed; requirements for additional capital; environmental risks; and general business and economic conditions. Often, but not always, forward-looking statements can be identified by the use of words such as “plans”, “expects”, “is expected”, “budget”, “scheduled”, “estimates”, “continues”, “forecasts”, “projects”, “predicts”, “intends”, “anticipates” or “believes”, or variations of, or the negatives of, such words and phrases, or statements that certain actions, events or results “may”, “could”, “would”, “should”, “might” or “will” be taken, occur or be achieved. Forward-looking statements involve known and unknown risks, uncertainties and other factors which may cause the Company and New Gold’s actual results, performance or achievements to be materially different from any of its future results, performance or achievements expressed or implied by forward-looking statements. These risks, uncertainties and other factors include, but are not limited to, the assumptions underlying the feasibility study not being realized, decrease of future gold prices, cost of labour, supplies, fuel and equipment rising, the availability of financing on attractive terms, actual results of current exploration, changes in project parameters, exchange rate fluctuations, delays and costs inherent to consulting and accommodating rights of First Nations, title risks, regulatory risks and uncertainties with respect to obtaining necessary surface rights and permits or delays in obtaining same, and other risks involved in the gold exploration and development industry, as well as those risk factors discussed in the section entitled “Description of Business-Risk Factors” in Rainy River Resources Ltd. latest Annual Information Form and its other SEDAR filings from time to time. Forward-looking statements are based on a number of assumptions which may prove to be incorrect, including, but not limited to, the availability of financing for the Company’s exploration and development activities; the timelines for the Company’s exploration and development activities on the Rainy River Gold property; the availability of certain consumables and services; assumptions made in mineral resource estimates, including geological interpretation grade, recovery rates, gold price assumption, and operational costs; and general business and economic conditions. All forward-looking statements herein are qualified by this cautionary statement. Accordingly, readers should not place undue reliance on forward-looking statements. The Company, New Gold and the authors of this report undertake no obligation to update publicly or otherwise revise any forward-looking statements whether as a result of new information or future events or otherwise, except as may be required by law.

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TABLE OF CONTENTS


1.	EXECUTIVE SUMMARY		1-1

1.1	Introduction		1-1

1.2	Contributors and Qualified Persons		1-2

1.3	Key Outcomes		1-3

1.4	Property Description		1-5

1.5	Accessibility, Climate, Local Resources, Infrastructure and Physiography		1-5

1.6	Project History		1-6

1.7	Geological Setting and Mineralization		1-7

1.8	Deposit Types		1-8

1.9	Exploration		1-8

1.10	Drilling		1-9

1.11	Sampling Method, Approach and Analyses		1-9

1.12	Data Verification		1-10

1.13	Metallurgical Testwork		1-11

	1.13.1	Historical Testwork	1-11

	1.13.2	Mineralogy	1-11

	1.13.3	Flowsheet Determination Testwork	1-12

	1.13.4	Comminution Tests	1-12

	1.13.5	Gravity Separation	1-13

	1.13.6	Cyanide Leaching	1-14

	1.13.7	Cyanide Destruction Testwork	1 15

	1.13.8	Environmental Testwork	1-15

	1.13.9	Testwork Interpretation	1-15

1.14	Mineral Resource Estimate		1-16

1.15	Open Pit and Underground Mine Design		1-19

	1.15.1	Open Pit Mine	1-19

	1.15.2	Underground Mine	1-21

	1.15.3	Open Pit and Underground Reserves	1-22

1.16	Mining Methods		1-23

	1.16.1	Open Pit Operations	1-25

	1.16.2	Underground Operations	1-27

	1.16.3	Proposed Mine Plan	1-30

1.17	Process Plant		1-32

1.18	Project Infrastructure		1-35

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	1.18.1	Site Water Management	1-39

	1.18.2	Tailings Management Area	1-39

1.19	Market Studies and Contracts		1-39

1.20	Environmental and Permitting		1-40

1.21	Capital and Operating Costs		1-41

1.22	Operating Costs		1-43

	1.22.1	Cash Costs	1-47

1.23	Economic Analysis		1-49

1.24	Adjacent Properties		1-52

1.25	Other Relevant Data and Information		1-53

1.26	Conclusions		1-54

1.27	Recommendations and Future Work Program		1-55

2.	INTRODUCTION		2-1

2.1	Scope of Study		2-1

2.2	Effective Dates and Declaration		2-2

2.3	Sources of Information		2-3

2.4	Terms of Reference		2-5

2.5	Site Visit		2-5

2.6	Acknowledgement		2-6

3.	RELIANCE ON OTHER EXPERTS		3-1

3.1	Report Responsibility and Qualified Persons		3-1

3.2	Other Study Contributors		3-4

4.	PROPERTY DESCRIPTION AND LOCATION		4-1

4.1	Mineral Tenure		4-1

4.2	Underlying Agreements		4-6

4.3	Environmental Considerations		4-6

4.4	Mining Rights in Ontario		4-9

5.	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY		5-1

5.1	Accessibility		5-1

5.2	Local Resources and Infrastructure		5-1

5.3	Climate		5-2

5.4	Physiography		5-2

6.	HISTORY		6-1

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6.1	Previous Exploration Work		6-1

	6.1.1	Period 1967 to 1989 by Various Companies	6-1

	6.1.2	Period 1990 to 2004 by Nuinsco	6-2

	6.1.3	Previous Mineral Resource Estimates	6-3

7.	GEOLOGICAL SETTING AND MINERALIZATION		7-1

7.1	Regional Geology		7-1

7.2	Property Geology		7-4

	7.2.1	Lithology	7-7

	7.2.2	Structural Geology	7-9

7.3	Mineralization		7-21

	7.3.1	Auriferous Sulphide and Quartz-Sulphide Stringers and Veins in Felsic Quartz-Phyric Rocks	7-22

	7.3.2	Deformed Quartz-Ankerite-Pyrite Shear Veins in Mafic Volcanic Rocks	7-26

	7.3.3	Silver-Rich Deformed Sulphide-Quartz Veins within Tuffaceous Rocks	7-26

	7.3.4	Nickel-Copper-PGE Mineralization	7-27

8.	DEPOSIT TYPES		8-1

9.	EXPLORATION		9-1

9.1	Period 1967-1989		9-1

9.2	Nuinsco Exploration Work (1990-2004)		9-1

9.3	Rainy River Exploration Work (2005-2012)		9-1

10.	DRILLING		10-1

10.1	Drilling from 2004 to 2011		10-1

	10.1.1	Introduction	10-1

	10.1.2	Drilling Procedures	10-3

10.2	Drilling Pattern and Density		10-4

10.3	SRK Comments		10-4

11.	SAMPLE PREPARATION, ANALYSES, AND SECURITY		11-1

11.1	Sample Preparation and Analyses		11-2

	11.1.1	Nuinsco Samples	11-2

	11.1.2	Rainy River Samples	11-2

	11.1.3	Metallurgical Testing	11-7

11.2	Specific Gravity Data		11-7

11.3	Quality Assurance and Quality Control Programs		11-9

11.4	SRK Comments		11-10

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project


12.	DATA VERIFICATION		12-1

12.1	Verification of Nuinsco Data		12-1

12.2	Verifications by Rainy River		12-1

12.3	Verifications by SRK		12-2

	12.3.1	Site Visit	12-2

	12.3.2	Verifications of Analytical Quality Control Data	12-2

	12.3.3	Verification of Electronic Data	12-4

13.	MINERAL PROCESSING AND METALLURGICAL TESTING		13-1

13.1	Historical Metallurgy		13-1

	13.1.1	Sample Selection	13-1

	13.1.2	Historical Testwork	13-1

13.2	Composite and Sample Selection		13-2

	13.2.1	Composite Selection	13-2

	13.2.2	Variability Sample Selection	13-3

13.3	Mineralogy		13-4

13.4	Flowsheet Selection Testwork		13-5

	13.4.1	Flotation Option	13-6

	13.4.2	Gravity and Gravity Tailings Whole Rock Leach	13-14

	13.4.3	Flowsheet Selection	13 16

13.5	Comminution Tests		13-16

	13.5.1	Crusher Work Index	13-16

	13.5.2	Unconfined Compressive Strength	13-17

	13.5.3	JK Drop Weight and SAG Mill Comminution	13-17

	13.5.4	SAGDesign	13-20

	13.5.5	Bond Work Index	13-22

	13.5.6	ModBond and Axb	13-24

	13.5.7	Bond Abrasion Index	13-25

	13.5.8	Comparison of Grinding Results from 2011 to 2012	13-26

	13.5.9	Grinding Circuit Design	13-27

13.6	Gravity Separation		13-29

	13.6.1	Gravity Recoverable Gold	13-29

	13.6.2	Variability Gravity Testwork	13-30

13.7	Heap Leaching		13-32

13.8	Cyanide Leaching		13-33

	13.8.1	Gravity Tailings Leaching	13-33

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	13.8.2	Cyanide Concentration	13-36

	13.8.3	Pre-Conditioning	13-37

	13.8.4	Oxygen (O₂) vs. Air	13-38

	13.8.5	Lead Nitrate Addition	13-39

	13.8.6	Grade Recovery Variability Tests	13-40

	13.8.7	Diagnostic Leach Testwork	13-44

13.9	Cyanide Destruction Testwork		13-45

13.10	Carbon-in-Pulp Modelling		13-45

13.11	Thickener Sizing Testwork		13-47

	13.11.1	Flocculant Screening	13-47

	13.11.2	Sedimentation Testwork	13-48

13.12	Rheology		13-49

13.13	Linear Screen Sizing Testwork		13-50

13.14	Environmental Testwork		13-51

13.15	Gold and Silver Recovery Curves		13-52

13.16	Testwork Interpretation		13-54

14.	MINERAL RESOURCE ESTIMATION		14-1

14.1	Introduction		14-1

14.2	Resource Estimation Procedures		14-2

14.3	Resource Database		14-2

14.4	Solid Body Modelling		14-4

	14.4.1	Introduction	14-4

	14.1.1	The ODM/17 Zone	14-6

	14.1.2	The 433, HS and New Zones	14-7

14.5	Compositing		14-10

14.6	Evaluation of Outliers		14-10

14.7	Statistical Analysis and Variography		14-13

	14.7.1	Statistical Analysis	14-13

	14.7.2	Variography	14-23

14.8	Block Model and Grade Estimation		14-29

	14.8.1	Block Model Definition	14-29

	14.8.2	Grade and Specific Gravity Estimation	14-29

14.9	Model Validation and Sensitivity		14-35

14.10	Mineral Resource Classification		14-36

14.11	Mineral Resource Statement		14-38

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14.12	Grade Sensitivity Analysis		14-46

14.13	Previous Mineral Resource Estimates		14-51

15.	MINERAL RESERVE ESTIMATE		15-1

15.1	Open Pit Mining		15-1

	15.1.1	Resource Block Model	15-1

	15.1.2	Open Pit Optimization	15-5

	15.1.3	Detailed Mine Design	15-11

	15.1.4	In-Pit Dilution and Mine Recovery	15-19

	15.1.5	Open Pit Material Inventory	15-19

15.2	Underground Mining		15-20

	15.2.1	Underground Mineral Reserves	15-20

	15.2.2	Underground Mineral Reserve Estimate	15-22

	15.2.3	Dilution and Mine Recovery	15-23

	15.2.4	Mining Recovery	15-25

	15.2.5	Underground Material Inventory	15-26

15.3	Open Pit and Underground Mineral Reserves		15-27

16.	MINING METHODS		16-1

16.1	Introduction		16-1

16.2	Open Pit Mining Methods		16-2

	16.2.1	Open Pit and Stockpiles Geotechnical Designs	16-2

	16.2.2	Open Pit Mine Planning	16-5

	16.2.3	Open Pit Mine Production Schedule	16-6

	16.2.4	Material Management	16-13

	16.2.5	Open Pit Mine Equipment and Operations	16-19

	16.2.6	Open Pit Mine Personnel Requirements	16-31

16.3	Underground Mining Methods		16-34

	16.3.1	Underground Geomechanical Design	16-36

	16.3.2	Longhole Open Stoping	16-41

	16.3.3	Cut and Fill	16-44

	16.3.4	Backfill	16-46

	16.3.5	Underground Mine Design and Schedule	16-47

	16.3.6	Definition Drilling and Sampling	16-50

	16.3.7	Underground Mine Services	16-50

	16.3.8	Underground Mine Schedule	16-57

16.4	Combined Production Schedule		16-62

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17.	RECOVERY METHODS		17-1

17.1	Proposed Process Flowsheet		17-1

17.2	Process Design Criteria		17-3

17.3	Process and Plant Facilities Description and Design Characteristics		17-4

	17.3.1	Primary Crushing	17-6

	17.3.2	Crushed Rock Handling and Storage	17-7

	17.3.3	Processing Plant and Tailings Handling	17-7

	17.3.4	Primary and Secondary Grinding	17-9

	17.3.5	Gravity Circuit	17-10

	17.3.6	Cyanide Leaching Circuit	17-10

	17.3.7	Carbon-in-Pulp Circuit, Carbon Stripping and Reactivation	17-11

	17.3.8	Tailings Management and Cyanide Destruction	17-12

	17.3.9	Refining Area and Gold Room	17-13

	17.3.10	Reagent Areas	17-13

	17.3.11	Control Room and Maintenance Shop	17-14

	17.3.12	Offices and Change House	17-14

	17.3.13	Metallurgical Laboratory	17-14

17.4	Energy, Water and Consumable Requirements		17-14

	17.4.1	Energy	17-14

	17.4.2	Water	17-15

	17.4.3	Consumables	17-17

18.	PROJECT INFRASTRUCTURE		18-1

18.1	General Site Works		18-1

	18.1.1	Primary Site and Access Roads	18-2

	18.1.2	Mine Haul Roads	18-2

18.2	Geotechnical		18-3

18.3	Mine Services Facilities		18-4

18.4	General Offices and Assay Laboratory		18-5

	18.4.1	Main Administration Building	18-5

	18.4.2	Mine Office and Dry	18-5

	18.4.3	Plant Office	18-6

18.5	Parking Area		18-6

18.6	Assay Lab		18-6

18.7	Fuel Storage and Dispensing		18-6

18.8	Explosive Plant and Storage		18-7

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18.9	Electrical and Communication		18-7

	18.9.1	Tie-Point Switching Station, Power Line, Main Substation, and Site Electrical Distribution	18-7

	18.9.2	Emergency Power	18-9

	18.9.3	Communication	18-11

18.10	Tailings Management		18-11

	18.10.1	Tailings Deposition Plan	18-12

	18.10.2	Dam Design	18-13

	18.10.3	Construction and Operational Considerations	18-14

	18.10.4	Site Water Management	18-17

	18.10.5	Water Management Structures	18-18

	18.10.6	Runoff and Seepage Collection	18-20

	18.10.7	Process Plant Water Supply – Preparations for Start-up	18-23

	18.10.8	Process Plant Water Supply – Operations	18-24

19.	MARKET STUDIES AND CONTRACTS		19-1

19.1	Market Studies		19-1

19.2	Contracts		19-1

20.	ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT		20-1

20.1	General Approach		20-1

20.2	Consultation Activities		20-1

	20.2.1	Community and Government Communications	20-1

	20.2.2	Aboriginal Communications	20-2

20.3	Environmental Studies		20-5

	20.3.1	Overview	20-5

	20.3.2	Climate, Air Quality and Sound	20-7

	20.3.3	Physiography, Soils and Geology	20-8

	20.3.4	Hydrology and Hydrogeology	20-11

	20.3.5	Surface Water, Sediment and Groundwater Quality	20-12

	20.3.6	Biological Environment - Existing Conditions	20-14

	20.3.7	Human Environment	20-18

20.4	Environmental Sensitivities		20-20

20.5	Regulatory Context		20-20

	20.5.1	Current Regulatory Status	20-20

	20.5.2	Environmental Approvals Required for Proposed Operations	20-21

20.6	Preliminary Environmental Impacts		20-26

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20.7	Preliminary Reclamation Plan		20-28

21.	CAPITAL AND OPERATING COSTS		21-1

21.1	Capital Cost Summary		21-2

21.2	Overhead Power Line		21-3

21.3	Highway 600 Replacement		21-3

21.4	Open Pit Mining Capital Cost Estimate		21-3

21.5	Underground Mining Capital Costs		21-5

21.6	Process Plant and Site Infrastructure Capital Costs		21-14

21.7	Process Plant Capital Costs		21-14

21.8	Tailings and Water Management Capital Costs		21-15

21.9	Rehabilitation and Mine Closure Costs		21-15

21.10	Direct Cost - Basis of Estimate		21-15

21.11	Indirect Cost - Basis of Estimate		21-19

21.12	Process Plant and Infrastructure Sustaining Costs		21-22

21.13	Exclusions		21-23

21.14	Assumptions		21-23

21.15	Operating Costs Summary		21-23

	21.15.1	Power and Fuel Costs	21-25

	21.15.2	Open Pit Mining Operating Costs	21-26

	21.15.3	Underground Mining Operating Costs	21-30

	21.15.4	Process Plant Operating Costs	21-32

	21.15.5	General and Administrative Costs	21-34

	21.15.6	Royalties	21-36

	21.15.7	Transportation and Refining	21-36

22.	ECONOMIC ANALYSIS		22-1

22.1	Assumptions and Basis		22-1

22.2	Royalties		22-3

22.3	Taxation		22-3

22.4	Financial Analysis Summary		22-4

22.5	Financial Model Sensitivity Analysis		22-7

23.	ADJACENT PROPERTIES		23-1

24.	OTHER RELEVANT DATA AND INFORMATION		24-1

24.1	Project Execution		24-1

24.2	Health, Safety, Environmental and Security		24-1

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24.3	Hazardous Waste Management		24-1

24.4	Execution Strategy		24-2

24.5	Management Procedures		24-2

24.6	Project Scheduling		24-4

24.7	Procurement and Contracts		24-6

24.8	Site Development		24-6

24.9	Construction		24-6

	24.9.1	Construction Management Responsibilities	24-6

	24.9.2	Construction Power	24-7

	24.9.3	Construction Labour Requirement	24-7

24.10	Process Facilities		24-8

	24.10.1	Critical Path and Installation Methodology	24-8

24.11	Tailings Management Area Earthworks		24-9

24.12	Commissioning		24-9

24.13	Mechanical Completion		24-10

24.14	Risk Management		24-10

25.	INTERPRETATION AND CONCLUSIONS		25-1

25.1	Sampling Method, Approach and Analyses		25-1

25.2	Data Verification		25-1

25.3	Mineral Resources		25-2

25.4	Sampling Preparation, Analysis and Security		25-4

25.5	Mining Methods and Reserves		25-4

25.6	Metallurgy and Processing		25-5

25.7	Infrastructure		25-6

25.8	Environmental Permitting		25-8

25.9	Financial Analysis		25-8

25.10	Conclusion		25-9

26.	RECOMMENDATIONS		26-1

26.1	Proposed 2013 Work Program		26-1

26.2	Further Recommendations		26-2

27.	REFERENCES		27-1

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APPENDICES

APPENDIX A – Rainy River Resources Title Opinion (As of March 13, 2013)

APPENDIX B – Rainy River Resources Patented, Leasehold and Mining Claims (March 13, 2013)

APPENDIX C – Nuinsco Exploration Activities (1993 – 2002)

APPENDIX D – Analytical Quality Control Data and Relative Precision Charts (Dec 2011 – Jul 2012)

APPENDIX E – Domain Variagrams

APPENDIX F – Geological Plan Map and Cross-Sections

APPENDIX G – Quantile-Quantile Plots - Block Data vs. Declustered Capped Composite Data

APPENDIX H – Rainy River Gold Project Site Plans

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List of Tables


Table 1-1:	Feasibility Study Qualified Persons	1-2

Table 1-2:	Key Outcomes – Gold Production	1-3

Table 1-3:	Key Outcomes – Silver Production	1-4

Table 1-4:	Key Outcomes – Capital Expenses and Financial Results	1-4

Table 1-5:	Mineral Resource Statement, Rainy River Gold Project, Ontario,	1-18

Table 1-6:	Pit Optimization Parameters	1-20

Table 1-7:	Open Pit Mine Design Parameters	1-21

Table 1-8:	Open Pit and Underground Proven and Probable Mineral Reserves (April 10, 2013)	1-23

Table 1-9:	Open Pit Mine Schedule	1-26

Table 1-10:	Underground Mine Schedule	1-29

Table 1-11:	Total Milled OP and UG	1-31

Table 1-12:	Overall Project Capital Cost Summary	1-42

Table 1-13:	Key Project Operating Costs	1-44

Table 1-14:	Underground Mining Operating Costs	1-46

Table 1-15:	Average Gold Cash Costs	1-48

Table 1-16:	Financial Analysis Summary	1-50

Table 1-17:	Rainy River Project Development Activities	1-53

Table 1-18:	Proven and Probable Reserves (April 10, 2013)	1-54

Table 1-19:	Budget for 2013	1-55

Table 2-1:	Major Study Contributors	2-2

Table 3-1:	Qualified Persons and Areas of Report Responsibility	3-2

Table 6-1:	Mineral Resource Statement* for the Rainy River Gold Project, Ontario, Mackie et al., December 23, 2003	6-3

Table 6-2:	Mineral Resource Statement for the Rainy River Gold Project, Ontario, Caracle Creek International Consulting Inc., April 30, 2008	6-4

Table 6-3:	Mineral Resource Statement* for the Rainy River Gold Project, Ontario, SRK Consulting (Canada) Inc., April 28, 2009	6-4

Table 6-4:	Mineral Resource Statement*, Rainy River Gold Project, Ontario, SRK Consulting (Canada) Inc., February 26, 2010	6-6

Table 6-5:	Mineral Resource Statement*, Rainy River Gold Project, Ontario, SRK Consulting (Canada) Inc., February 24, 2011	6-7

Table 6-6:	Mineral Resource Statement*, Rainy River Gold Project, Ontario, SRK Consulting (Canada) Inc., June, 29 2011	6-8

Table 6-7:	Mineral Resource Statement*, Rainy River Gold Project, Ontario, SRK Consulting (Canada) Inc., February, 24 2012	6-9

Table 9-1:	Summary of Exploration Work by Rainy River on the Rainy River Gold Project between 2005 and 2012	9-5

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Table 10-1:	Core Drilling Completed on the Rainy River Gold Project (1994-2012)	10-1

Table 11-1:	Specific Gravity Assigned to Gold Zones	11-9

Table 12-1:	Summary of Analytical Quality Control Data Produced between December 2011 and July 2012	12-3

Table 13-1:	Master Composite Proportions	13-1

Table 13-2:	Weight Percentages by Zone for Initial Pit and RLOM Composites and Overall Pit	13-3

Table 13-3:	Rougher Flotation Results	13-7

Table 13-4:	Cleaner Flotation Results	13-8

Table 13-5:	Flotation Concentrate Leaching Results (Gold Assays)	13-9

Table 13-6:	Flotation Concentrate Leaching Results (Silver Assays)	13-9

Table 13-7:	Flotation Tailings Leaching Results	13-10

Table 13-8:	IsaMill Testwork Results	13-11

Table 13-9:	Stirred Media Detritor Testwork Results	13-13

Table 13-10:	Gravity Tailings Leach Results (Gold)	13-15

Table 13-11:	Gravity Tailings Leach Results (Silver)	13-15

Table 13-12:	Crusher Work Index Results	13-16

Table 13-13:	JK Drop Weight (DWT) and Corresponding SMC Results	13-18

Table 13-14:	SAG Mill Comminution SMC and M_iaValues	13-19

Table 13-15:	SAGDesign Testwork Results	13-20

Table 13-16:	Comparison between SAGDesign and JK DWT Results	13-21

Table 13-17:	Bond and ModBond Results	13 23

Table 13-18:	Abrasion Index Results	13-25

Table 13-19:	SMC (Axb) Comparison of 2011 vs. 2012	13-26

Table 13-20:	SAG and Ball Mill Simulation Results	13-28

Table 13-21:	Gravity Recoverable Gold Results	13-30

Table 13-22:	Additional Gravity Tailings Leaching Results (Gold Assays)	13-33

Table 13-23:	Additional Gravity Tailings Leaching Results (Silver Assays)	13-34

Table 13-24:	Cyanide Concentration Testwork Results	13-36

Table 13-25:	Pre-Conditioning vs. No Pre-Conditioning Testwork Results	13-38

Table 13-26:	O₂vs. Air and Lead Nitrate Addition Testwork Results	13-39

Table 13-27:	Gold Leaching Variability Testwork Average Results	13-41

Table 13-28:	Silver Leaching Variability Testwork Average Results	13-41

Table 13-29:	Cyanide Destruction Testwork Results	13-45

Table 13-30:	Flocculant Description	13-47

Table 13-31:	Sedimentation Testwork Results	13-48

Table 13-32:	Linear Screen Sizing Testwork Results	13-51

Table 13-33:	Summary of Geochemical Environmental Testing	13-52

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Table 13-34:	Residue and Gravity Recovery Curves	13-53

Table 13-35:	Gold and Silver Recoveries vs. Head Grade	13-54

Table 13-36:	Average Yearly Gold and Silver Recoveries	13-56

Table 14-1:	Rock Codes in the Rainy River Gold Project Block Model	14-4

Table 14-2:	Summary of Metal Capping Levels Applied to Each Resource Domain	14-11

Table 14-3:	Basic Statistics for Gold Composites for All Resource Domains	14-14

Table 14-4:	Base Statistics for Capped Gold Composites for All Resource Domains	14-15

Table 14-5:	Basic Statistics for Silver Composites for All Resource Domains	14-16

Table 14-6:	Base Statistics for Capped Silver Composites for All Resource Domains	14-17

Table 14-7:	Basic Statistics of Composites for Domain 200 (Zone 34)	14-18

Table 14-8:	Basic Statistics for Calcium Uncapped Composites for All Resource Domains	14-19

Table 14-9:	Basic Statistics for Calcium Capped Composites for All Resource Domains	14-20

Table 14-10:	Basic Statistics for Sulphur Uncapped Composites for All Resource Domains	14-21

Table 14-11:	Basic Statistics for Sulphur Capped Composites for All Resource Domains	14-22

Table 14-12:	Modeled Gold Variogram Parameters for All Resource Domains Grade Interpolation	14-25

Table 14-13:	Modeled Silver Variogram Parameters for All Resource Domains Grade Interpolation	14-26

Table 14-14:	Modeled Calcium Variogram Parameters for All Resource Domains Grade Interpolation	14-27

Table 14-15:	Modeled Sulphur Variogram Considered for All Resource Domains Grade Interpolation	14-28

Table 14-16:	Rainy River Gold Project Block Model Parameters	14-29

Table 14-17:	Resource Estimation Parameters	14-31

Table 14-18:	Search Neighbourhoods Used for Gold and Silver Estimation	14-32

Table 14-19:	Search Neighbourhoods Used for Calcium	14-33

Table 14-20:	Search Neighbourhoods Used for Sulphur	14-34

Table 14-21:	Search Parameters Used for Grade Estimation in Zone 34	14-35

Table 14-22:	Search Parameters Used to Code the Measured Blocks	14-37

Table 14-23:	Conceptual Assumptions Considered for Open Pit Resource Reporting	14-39

Table 14-24:	Conceptual Assumptions Considered for Underground Resource Reporting	14-39

Table 14-25:	Consolidated Mineral Resource Statement*, Rainy River Gold Project, Ontario, SRK Consulting (Canada) Inc., October 10, 2012	14-43

Table 14-26:	Mineral Resources* for the 34 Zone (Domain 200), Rainy River Gold Project, Ontario, SRK Consulting (Canada) Inc., October 10, 2012	14-44

Table 14-27:	Mineral Resources* for the Silver Zone (Domain 901), Rainy River Gold Project, SRK Consulting (Canada) Inc., October 10, 2012	14-44

Table 14-28:	Open Pit Mineral Resources*, Rainy River Gold Project, Ontario, SRK Consulting (Canada) Inc., October 10, 2012	14-45

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Table 14-29:	Underground Mineral Resources*, Rainy River Gold Project, Ontario, SRK Consulting (Canada) Inc., October 10, 2012	14-46

Table 14-30:	Global Block Model Quantities and Grade Estimates* at Various Cut-Off Grades	14-47

Table 14-31:	Block Model Quantities and Grade Estimates* at Selective Cut-off Grades Potential Open Pit Mining Material	14-49

Table 14-32:	Block Model Quantities and Grade Estimates* at Selected Cut-off Grades – Potential Underground Mining Material	14-50

Table 14-33:	Comparison of February 2012 and October 2012 Mineral Resource Statements	14-52

Table 15-1:	Au Recovery by Grade and Ore Type	15-4

Table 15-2:	Pit Optimization Parameters	15-6

Table 15-3:	Selected Pits at Various Cut-Off Grades	15-10

Table 15-4:	Detailed Open Pit Mine Design Parameters	15-11

Table 15-5:	Estimation of In-pit Dilution and Mine Recovery	15-19

Table 15-6:	Combined Open Pit and Underground Mineral Resource Estimated by SRK and Underground Mineral Resource Estimate	15-21

Table 15-7:	MSO Input Parameters	15-22

Table 15-8:	Total Dilution Estimates from Waste Rock by Mining Method	15-24

Table 15-9:	Total Dilution Estimate from Mucking Backfill by Mining Method	15 24

Table 15-10	Total Dilution Estimate from Cemented Rock Fill Wall Sloughing by Mining Method	15-25

Table 15-11:	Total Dilution Estimate by Mining Method	15-25

Table 15-12:	Mining Recovery Factors used to Estimate the Mineral Reserves	15-26

Table 15-13:	Total Underground Mineral Reserves by Mining Method (COG of 3.5 g/t Au eq)	15-27

Table 15-14:	Total Underground Mineral Reserves by Category (COG of 3.5 g/t Au eq)	15-27

Table 15-15:	Open Pit and Underground Proven and Probable Mineral Reserves (April 10, 2013)	15-28

Table 16-1:	Recommended Overall Slope Geometry by Sector	16-3

Table 16-2:	Waste Rock and Low-Grade Ore Stockpile Design Parameters	16-16

Table 16-3:	NPAG Pile Summary	16-16

Table 16-4:	PAG Pile Summary	16-16

Table 16-5:	Low-Grade Ore Stockpile Summary	16-17

Table 16-6:	Overburden Pile Design Parameters	16 18

Table 16-7:	Overburden Pile Design Summary	16-18

Table 16-8:	Operating Shift Parameters	16-21

Table 16-9:	Equipment and Worker Operating Time	16-21

Table 16-10:	Major Equipment Availability and Utilization	16-22

Table 16-11:	Drill and Blast Specifications	16-24

Table 16-12:	Trucks Speeds and Fuel Consumptions (Loaded and Empty)	16-26

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Table 16-13:	Annual Open Pit Mine Equipment Requirements	16 30

Table 16-14:	Annual Hourly Personnel Requirements	16-32

Table 16-15:	Salaried Open Pit Personnel Requirements	16-33

Table 16-16:	Preliminary Longhole Open Stoping Design Parameters	16-36

Table 16-17:	Preliminary Cut and Fill Design Parameters	16-36

Table 16-18:	Modified Stability Graph for ODM West Zone Above 500 m Indicating Stability of Stope Surfaces Based on Potential Design Limits and Actual Final Stope Dimensions	16-38

Table 16-19:	Ground Support Recommendations	16-39

Table 16-20:	Design Details of the Mine Access Infrastructure	16-47

Table 16-21:	Design Details for the Level Access Infrastructure	16-48

Table 16-22:	Design Details for the Footwall Drifts and Level Access	16-49

Table 16-23:	Design Details for Longhole Open Stoping Drawpoint	16-49

Table 16-24:	Rainy River Ventilation Model – Peak Production Air Quantities	16-51

Table 16-25:	Rainy River Ventilation Model – Raise Details	16-52

Table 16-26:	Rainy River Ventilation Model – Fan Summary	16-53

Table 16-27:	Summary of Sump Details for the Rainy River Underground	16-55

Table 16-28:	Development Phase Equipment Requirements	16-58

Table 16-29:	Stoping Productivities Used in the Mine Schedule	16 61

Table 16-30:	Open Pit and Underground Mine Production Schedule	16-63

Table 17-1:	General Process Design Criteria	17-3

Table 17-2:	Process Plant Power Demand by Area	17-14

Table 17-3:	Process Plant Reagent Consumption	17-21

Table 17-4:	Grinding Media Consumptions by Mill Type	17-21

Table 18-1:	Mining Vehicle Dimensions	18-4

Table 18-2:	Mining Vehicle Repair Bay Specifications	18-4

Table 18-3:	Staff Requirements	18-5

Table 18-4:	Estimated Total Project Power Demand	18-8

Table 18-5:	Emergency Power Requirements	18 10

Table 18-6:	Proposed Emergency Generators	18-10

Table 18-7:	Tailings Dam Sizing	18-12

Table 18-8:	Construction Fill Materials	18-17

Table 18-9:	Water Availability from the Pinewood River below the McCallum Creek Inflow	18-23

Table 20-1:	Species at Risk Known to be Present in the Rainy River Gold Project Environs	20-18

Table 20-2:	Potential Federal Approvals	20-24

Table 20-3:	Potential Provincial Approvals	20-25

Table 20-4:	Preliminary Summary of Potential Environmental Effects	20-26

Table 20-5:	Preliminary Reclamation Approach	20-30

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Table 21-1:	Project Capital Cost Summary	21-2

Table 21-2:	Open Pit Mine Equipment Capital Cost	21-4

Table 21-3:	Mobile Equipment Capital Cost Estimate	21-6

Table 21-4:	Estimate of the Stationary and Auxiliary Equipment Capital Cost	21-8

Table 21-5:	Equipment Operating Cost used at the Rainy River Underground	21-9

Table 21-6:	Peak Staff Complement and Cost for the Rainy River Underground	21-10

Table 21-7:	Peak Hourly Complement and Cost for the Rainy River Underground	21-10

Table 21-8:	Cost of Major Consumable Items	21-11

Table 21-9:	Unit Company Development Costs	21-12

Table 21-10:	Contractor Development Cost	21-12

Table 21-11:	Rainy River Underground Capital Cost Summary	21 13

Table 21-12:	Infrastructure, Process Plant and TMA Pre-Production Costs	21-14

Table 21-13:	Labour Productivity Factors	21-18

Table 21-14:	Indirect Costs	21-19

Table 21-15:	Operating Cost Summary	21-24

Table 21-16:	Key Project Operating Costs	21-24

Table 21-17:	Open Pit Mine Operating Cost Breakdown	21-26

Table 21-18:	Salaried Mine Personnel Annual Cost (Open Pit)	21-28

Table 21-19:	Hourly Mine Personnel Annual Cost (Open Pit)	21-29

Table 21-20:	Breakdown of the Rainy River Underground Operating Cost by Activity	21-31

Table 21-21:	Rainy River Underground Yearly and Total Operating Cost (‘000)	21-31

Table 21-22:	Process Operating Costs	21-32

Table 21-23:	Yearly Average Processing Operating Cost Breakdown	21-34

Table 21-24:	Average General and Administrative Costs	21-36

Table 22-1:	Financial Model Criteria	22-2

Table 22-2:	Financial Analysis Summary (Pre-tax and After-tax)	22-5

Table 22-3:	Rainy River Financial Model Summary	22-6

Table 22-4:	Sensitivity Results for Metal Price and Exchange Rate Variations	22-8

Table 22-5:	Sensitivity Results for CAPEX, OPEX and Metal Recovery Variations (After-tax)	22-8

Table 25-1:	Mineral Resource Statement, Rainy River Gold Project, Ontario,	25-3

Table 25-2:	Comparison of February 2012 and October 2012 Mineral Resource Statements	25-4

Table 25-3:	Proven and Probable Mineral Reserves (Effective date April 10, 2013)	25-5

Table 26-1:	Budget for 2013	26-1

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List of Figures
Figure 1-1:	Initial Pit, Final Pit and Resource Pit Shell	1-27

Figure 1-2:	Isometric View of the Rainy River Underground Mine	1-30

Figure 1-3:	Annual Gold and Silver Production (koz.)	1-31

Figure 1-4:	Schematic Process Flowsheet	1-34

Figure 1-5:	General Site Layout	1-36

Figure 1-6:	Annual Operating Cash Costs (USD/oz. Au) with Silver Credit	1-48

Figure 1-7:	Life-of-Mine Cash Flow Projections	1-51

Figure 1-8:	After-Tax Net Present Value (NPV) Sensitivity Analysis at 5% Discount Rate	1-51

Figure 1-9:	After-Tax Internal Rate of Return (IRR) Sensitivity Analysis	1-52

Figure 4-1:	Location of Rainy River Gold Project (as of March 13, 2013)	4-3

Figure 4-2:	Land Tenure Map of the Rainy River Gold Project (as of March 13, 2013)	4-5

Figure 5-1:	Typical Landscape in the Rainy River Gold Project Area	5-3

Figure 7-1:	Regional Bedrock Geology of the Area West of Fort Frances	7-3

Figure 7-2:	Bedrock Geological Interpretation for the Area Surrounding the Rainy River Gold Project (from Rainy River, 2012)	7-5

Figure 7-3:	A North-South Geological Cross Section Across the Rainy River Deposit (Rainy River, 2012)	7-6

Figure 7-4:	Regional Structural Trends on the Rainy River Gold Project	7-11

Figure 7-5:	Structural Fabrics Affecting Rock Types	7-12

Figure 7-6:	Summary of Structural Data Collected by SRK in 2011	7-13

Figure 7-7:	Pressure Shadows Around Rigid Objects in Dacitic Rock from the ODM/17 Zone	7-14

Figure 7-8:	High Strain Zone at the Base of the ODM/17 Zone (SRK, 2011)	7-15

Figure 7-9:	Oblique 3D View of ODM/17 Zone Looking Down Plunge to the Southwest (SRK, 2011)	7-16

Figure 7-10:	Chlorite Carbonate Shear Zones (SRK, 2011)	7-17

Figure 7-11:	Brittle Ductile Shear Zones Identified in the ODM/17 Zone	7-18

Figure 7-12:	Evidence for Strike-Slip Kinematics of Late Brittle Faulting (SRK, 2011)	7-19

Figure 7-13:	Sulphide Mineralization Deformed by Folding in Core from the Rainy River Gold Project (SRK, 2011)	7-20

Figure 7-14:	Structural Control of the Plunge of the Gold Mineralization at the Rainy River Gold Project	7-21

Figure 7-15:	ODM/17 Zone Gold Mineralization (SRK, 2011)	7-23

Figure 7-16:	ODM/17 High-Grade Gold Mineralization (SRK, 2011)	7-24

Figure 7-17:	433 High-Grade Gold Mineralization (SRK, 2011)	7-25

Figure 7-18:	Higher Grade Gold Mineralization (SRK, 2011)	7-27

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Figure 8-1:	Idealized Sketch Showing Relative Timing of Auriferous Features and Illustrating Protracted Deformation Affecting Initial Gold-rich Volcanogenic Mineralization and Subsequently Overprinted by Mesothermal Gold Mineralization (SRK, 2011)	8-3

Figure 8-2:	Schematic Geological Setting and Hydrothermal Alteration Associated with Gold-rich Volcanogenic Hydrothermal Systems (after Hannington et al., 1999)	8-5

Figure 8-3:	Schematic Section and Hydrothermal Alteration Associated with the Rainy River Gold Project (Sparkes & Wartman, 2012)	8-6

Figure 9-1:	Section Showing the Footwall Silver Zone Below the previous PEA Starter Pit	9-3

Figure 10-1:	Core Drilling Data by Period (1994 to 2012)	10-2

Figure 10-2:	Drill Collar Map in Relation to Resource Domains and Conceptual Pit Outline	10-2

Figure 11-1:	Summary of Specific Gravity Composite Data. Top: All Resource Domains;	11-8

Figure 13-1:	Sample Locations for Comminution (Left) and Leaching (Right) Variability Testwork	13-3

Figure 13-2:	Sample Locations for Comminution (Left) and Leaching (Right)	13-4

Figure 13-3:	Specific Energy vs. Particle Size for IsaMill Test	13-12

Figure 13-4:	Specific Energy vs. Particle Size for SMD Test	13-14

Figure 13-5:	SMC Data Distribution with JK DWT Calibration Points	13-19

Figure 13-6:	Full Bond Ball Mill Work Indices vs. ModBond Work Indices (200 Mesh)	13-22

Figure 13-7:	Distribution of ModBond Indices (200 Mesh) by Zone	13-24

Figure 13-8:	ModBond Wi vs. Axb	13-25

Figure 13-9:	Gold Gravity Recovery vs. Head Grade	13-31

Figure 13-10:	Silver Gravity Recovery vs. Head Grade	13-31

Figure 13-11:	Heap Leach Gold Recovery Curve	13-32

Figure 13-12:	Gravity Tailings Leach Residue vs. Grind Size	13-34

Figure 13-13:	Cost and Revenue Analysis by Grind Size	13-35

Figure 13-14:	Gold Recovery vs. Time at Different NaCN Concentrations	13-37

Figure 13-15:	Gold Residue vs. Head Grade (Variability Tests)	13-43

Figure 13-16:	Silver Residue vs. Head Grade (Variability Tests)	13-43

Figure 13-17:	CIP Modeling Isotherms	13-46

Figure 13-18:	Initial Pit Composite Yield Stress vs. Solids Density	13-49

Figure 13-19:	Remaining-Life-of-Mine Composite Yield Stress vs. Solids Density	13-50

Figure 14-1:	Location of New Boreholes Drilled During the Period March to December 2011,	14-5

Figure 14-2:	Isometric View of the Rainy River Mineralization Wireframes Modelled by SRK with Borehole Data (View looking towards the west showing 2011 drilling only)	14-6

Figure 14-3:	Histogram Distribution of Raw Sample Lengths	14-10

Figure 14-4:	Cumulative Frequency Plot for Gold Composites	14-12

Figure 14-5:	Examples of Variogram Models for Rainy River Gold Deposit	14-24

Figure 14-6:	Cross-Section 425,775E Comparing Blocks Populated with Gold-Grades and Informing Data	14-36

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Figure 14-7:	Schematic Vertical Section Illustrating Criteria Considered for Preparing the Mineral Resource Statement for the Rainy River Gold Project (View Looking East)	14-41

Figure 14-8:	Rainy River Gold Project Global Grade Tonnage Curves	14-47

Figure 14-9:	Distribution of Open Pit Mineral Resources Relative to the Conceptual Pit Outline	14-51

Figure 15-1:	Isopach Mapping of Overburden Thickness	15-3

Figure 15-2:	Slope by Sector for Pit Optimization	15-7

Figure 15-3:	Surface Constraints for Pit Optimization	15-8

Figure 15-4:	Rainy River Theoretical Pit Shell (Plan View)	15-9

Figure 15-5:	Detailed Open Pit Mine Design (Plan View)	15-12

Figure 15-6:	Final Pit Design and LG Optimization – Isometric View	15-13

Figure 15-7:	Final Pit Design and LG Optimization – Elevation 300 masl	15-14

Figure 15-8:	Final Pit Design and LG Optimization – Elevation 150 masl	15-15

Figure 15-9:	Final Pit Design and LG Optimization – Elevation 0 masl	15-16

Figure 15-10:	Final Pit Design and LG Optimization – Cross Section (East 425 500, looking West)	15-17

Figure 15-11:	Final Pit Design and LG Optimization – Cross Section (East 425 800, looking West)	15-18

Figure 16-1:	Isometric View of the Rainy River Underground Mine	16-1

Figure 16-2:	Open Pit Design Zones & Recommendations (AMEC, 2013F)	16-4

Figure 16-3:	Starter Pit and Final Pit - 3D View	16-7

Figure 16-4:	Open Pit Mine Planning - 2016 Q2 (End of Pre-Production Period)	16-8

Figure 16-5:	Open Pit Mine Planning – 2017 (End of Year 2)	16-9

Figure 16-6:	Open Pit Mine Planning – 2019 (End of Year 4)	16-10

Figure 16-7:	Open Pit Mine Planning – 2021 (End of Year 6)	16-10

Figure 16-8:	Open Pit Mine Planning – 2023 (End of Year 8)	16-11

Figure 16-9:	Open Pit Material Movement over the Life-of-Mine	16-12

Figure 16-10:	Site Plan Showing West and East Dump Areas	16-14

Figure 16-11:	East Dump Area	16-14

Figure 16-12:	West Dump Area	16-15

Figure 16-13:	In-Pit Dumping Area	16-19

Figure 16-14:	Cycle Time Trend over LOM	16-27

Figure 16-15:	LOM Haul Truck Fleet	16-28

Figure 16-16:	Longitudinal Section of the Rainy River Resources Underground Mine	16-35

Figure 16-17:	View North West of the Underground Mine Geometry	16-37

Figure 16-18:	Empirical Estimation of Wall Slough (ELOS) for Varying HW dip Cases for the 4 Main Underground Design Zones above 500 m depth in which LHOS was applied	16-40

Figure 16-19:	Schematic of the Typical Progression of Transverse Longhole Open Stoping	16-42

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Figure 16-20:	Typical Longhole Open Stoping Drilling Layout (Ore Widths less than 8 m)	16-44

Figure 16-21:	General Longitudinal Section of the Cut and Fill Mining Method	16-45

Figure 16-22:	General Cut and Fill Ramp Arrangement	16-46

Figure 16-23:	Longitudinal Section of the Rainy River Underground Ventilation Layout	16-53

Figure 16-24:	Waste and Void Schedule from the Rainy River Underground (Golder 2013)	16-62

Figure 17-1:	Whole Rock Leach Process Schematic Diagram	17-2

Figure 17-2:	General Processing Area and Buildings Site Layout	17-5

Figure 17-3:	Process Plant General Arrangement Drawing (Including Primary Electrical Substation)	17-8

Figure 17-4:	Process Plant and Tailings Ponds Water Balance	17-16

Figure 18-1:	Typical Cross Section - TMA South Dam Section	18-16

Figure 18-2:	Typical Cross Section – TMA West and North Sections	18-16

Figure 21-1	- Cumulative Probability	21-21

Figure 21-2:	Annual Operating Cash Costs (USD/oz. Au) with Silver Credit	21-25

Figure 22-1:	Life-of-Mine Cash Flow Projection (Pre-tax and After-tax, discount rate: 5%)	22-7

Figure 22-2:	Sensitivity of the Net Present Value (After-tax) to Financial Variables	22-9

Figure 22-3:	Sensitivity of the Internal Rate of Return (After-tax) to Financial Variables	22-9

Figure 23-1:	Bayfield Ventures Corp. Properties	23-2

Figure 23-2:	Coventry Resources Properties	23-3

Figure 23-3:	King’s Bay Gold Menary Property	23-4

Figure 24-1:	Project Management Organization Chart	24-7

Figure 24-2:	Project Milestones	24-9

Figure 24-3:	Monthly Construction Manpower Graph	24-12

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LIST OF ABBREVIATIONS
$		Dollar sign
%		Percent sign
¢/kWh		Cent per kilowatt hour
°		Degrees
°C		Degrees Celsius
3D		3 Dimensional
a		Acres
AAS		Atomic Absorption Spectroscopy
Actlabs		Activation Laboratories
ALS		ALS Minerals Laboratories
AMT		Audio Magneto-Telluric
AI		Abrasion Index
APEGBC		Professional Engineers and Geoscientists of British Columbia
APEO		Association of Professional Engineers of Ontario
APGO		Association of Professional Geoscientists of Ontario
Au		Gold
bank		Bank cubic metre - volume in-situ
BFA		Bench Face Angles
BWI		Bond Ball Mill Work Index
CAD		Canadian Dollar
CAF		Cut and Fill
CAPEX		Capital Expense Estimate
CICC		Caracle Creek International Consulting Inc.
CIL		Carbon-in-Leach
CIM		Canadian Institute of Mining, Metallurgy and Petroleum
CIP		Carbon-in-Pulp
CLRA		Construction Labour Relations Association
CM		Construction Management
CMT		Corporate Minimum Tax
CRF		Cemented Rock Fill

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LIST OF ABBREVIATIONS
CSA		Canadian Standards Assocation
CWI		Crusher Work Index
D2		Second generation of deformation
D3		Third generation of deformation
D4		Fourth generation of deformation
DDH		Diamond Drill Hole
DGPS		Differential Global Positioning System
DWI		Drop Weight Index
DWT		Drop Weight Test
DXF		Drawing Interchange Format
E		East
EA		Environmental Assessment
EAC		Estimate at Completion
EDF		Environmental Design Flood
ELC		Extended Life Coolants
ELOS		Equivalent Linear Overbreak/Slough
EO		Enterprise Optimization
EP		Engineering and Procurement
EPCM		Engineering, Procurement and Construction Management
FAR		Fresh Air Raise
FFCS		Fort Frances Chiefs Secretariat
FOS		Factor of Safety
FS		Feasibility Study
FW		Footwall
g		Grams
g/t		Grams per tonne
gal.		Gallons
G&A		General and Administrative
GA		General Arrangement
GRG		Gravity Reconverable Gold

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LIST OF ABBREVIATIONS
GSI		Geological Strength Index
h		Hour
ha		Hectare
HAZOP		Hazard and Operability
HBED		Hudson’s Bay Exploration and Development
HP		Horsepower
HPDE		High-density polyethylene
HQ		Drill core size (6.4 cm diameter)
HS		High Strain (zone)
HSE		Health Safety and Environmental
HVAC		Heating, Ventilation and Air-Conditioning
HW		Hangingwall
IESO		Independent Electricity System Operator
IFRS		International Financial Reporting Standards
INCO		International Nickel Corporation of Canada
IRR		Internal Rate of Return
IT		Information Technology
JEF		Job Efficiency Factor
KCB		Klohn Crippen Berger Ltd.
kPa		Kilopascal
kV		Kilovolt
kW		kilowatt
kWh		kilowatt-hour
LG		Lerchs-Grossmann
LH		Longhole
LHD		Longhole Drill
LHOS		Longhole Open Stoping
LIMS		Local Information Management System
LOM		Life-of-Mine
M		Million

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LIST OF ABBREVIATIONS
m/h		Metres/hour
Ma		Million years
Masl		Meters above sea level
Mm³		Million cubic metres
MMBtu/h		1 Million British Thermal Units Per Hour
MMI		Mobile Metal Ion
MNO		Métis Nation of Ontario
MNDM		Ministry of Northern Development and Mines
MNR		Ministry of Natural Resources
MOE		Ministry of Environment
MOT		Ministry of Transportation of Ontario
MOU		Memorandum of Understanding
MPC		Mine Power Centre
MRP		Mine Rock Pond
MSE		Mechanically Stabilized Earth
MSO		Mineable Shape Optimizer
Mt		Million tonnes
Mtpa		Million tonnes per annum
MTO		Material Take-Offs
MW		Megawatt
N		North
NAG		Non-Acid Generating
Nb		number
NOH		Net Operating Hours
NP		Acid Neutralizing Capacity
NPAG		Not Potentially Acid-Generating
NPI		Net Profits Interest
NPR		Neutralization Potential Ratio
NPV		Net Present Value
NQ		Drill core size (4.8 cm diameter)

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LIST OF ABBREVIATIONS
NSR		Net Smelter Return
OB		Overburden
OIQ		Ordre des ingénieurs du Québec
OGS		Ontario Geological Survey
OMC		Oreway Mineral Consultants
OMT		Ontario Mining Tax
OP		Open pit
OPEX		Operational Expenditure
OSA		Overall Slope Angles
Owner		Rainy River Resources Ltd.
OZ		Ore Zones
oz.		ounce
oz./t		ounce per tonne
PA		Participation Agreement
PAAC		Participation Agreement Advisory Committee
PAG		Potential Acid-Generating
PEA		Preliminary Economic Assessment
PEO		Professional Engineers Ontario
PEP		Project Execution Plan
PGA		Peak Ground Acceleration
pH		percentage hydrogen
ppb		part per billion
ppm		part per million
Q1		First Quarter
Q2		Second Quarter
Q3		Third Quarter
Q4		Fourth Quarter
QA/QC		Quality Assurance/Quality Control
RF		Unconsolidated Rock Fill
RLOM		Remaining-Life-of-Mine

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LIST OF ABBREVIATIONS
RPM		Revolutions Per Minute
RQD		Rock Quality Designation
RRGB		Rainy River Green Belt
RRGP		Rainy River Gold Project
RRR		Rainy River Resources Ltd.
RRU		Rainy River Underground
RWI		Bond Rod Mill Work Index
S		South
SAG		Semi-Autogenous
SEDAR		System for Electronic Document Analysis and Retrieval
SG		Specific Gravity
SMC		SAG Mill Comminution
SMD		Stirred Mill Detritor
SP		Stockpile Pond
t/h		Tonnes per hour
tpa		Tonnes per annum
tpd		Tonnes per day
TMA		Tailings Management Area
TMAP		Tailings Management Area Pond
TSF		Tailings Storage Facility
TSX		Toronto Stock Exchange
UCS		Uniaxial Compressive Strength
UG		Underground
USD		United States Dollar
UTM		Universal Transverse Mercator
VMS		Volcanogenic Massive Sulphide
VOIP		Voice Over Internet Protocol
VPSA		Vacuum Pressure Swing Adsorption
W		West
WBS		Work Breakdown Structure

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LIST OF ABBREVIATIONS
WCP		West Creek Pond
WMP		Water Management Pond
XRD		X-ray diffraction
X		X coordinate (E-W)
µm		Microns
Y		Y coordinate (N-S)
Z		Z coordinate (depth or elevation)

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

1.	EXECUTIVE SUMMARY

1.1

Introduction

The Rainy River Gold Project (sometimes referred to as the “Project”) is an advanced gold exploration project located 50 km northwest of Fort Frances, in northwestern Ontario. Rainy River Resources Ltd. (“RRR”, “Rainy River” or “the Company”) holds a 100% interest in the Project.

The Feasibility Study of the Project discussed in this Technical Report, has been prepared at the request of Rainy River and filed by Rainy River on SEDAR (www.sedar.com). Pursuant to a takeover bid commenced on June 18, 2013, New Gold Inc. (“New Gold”) has acqured majority ownership of RRR. RRR continues to exist as a separate legal entity from New Gold. New Gold has requested that BBA readdress this Technical Report to New Gold in order to support its own disclosure. No changes have been made to this Technical Report beyond addressing it to New Gold, inserting references to New Gold where appropriate and re-dating it, as well as formatting changes and minor typographical corrections.

The Feasibility Study was prepared and compiled by BBA Inc. (“BBA”) in a collaborative effort with Rainy River, together with a number of specialized consultants. This Technical Report was prepared according to the guidelines set out under the requirements of National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”).

The main intent of the Feasibility Study is to provide a technical and economic review of the potential mining operations based on the Company’s most recent mineral resource estimate and other project definition activities such as metallurgical testwork. The mineral resources (including inferred resources) used in the Feasibility Study are based on updated information issued in a press release by Rainy River on October 10, 2012, entitled “Rainy River Resources Announces 6.2 Million Ounces of Gold in Measured and Indicated Resources, and 2.3 Million Ounces of Gold in Inferred Resources”. The Feasibility Study is based on the development of a combined open pit and underground mining operation, feeding a process plant to recover gold and silver mineralization. The process plant will have a capacity of 21,000 tonnes per day.

All monetary units in the Report are in Canadian dollars (“CAD”), unless otherwise specified. Costs are based on fourth quarter (“Q4”) 2012 dollars.

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1.2

Contributors and Qualified Persons

This Technical Report has been prepared for Rainy River based on work prepared by a number of independent consultants. A summary of the Qualified Persons and Study contributors, their area of responsibility and site visit dates are listed in Table 1-1.

Table 1-1: Feasibility Study Qualified Persons


Consulting Firm or Entity	Area of Responsibility	Qualified Person(s)	Site Visits

SRK Consulting (Canada) Inc.	Geological modelling and resource definition.	Glen Cole Dorota El-Rassi	April 30-May 2, 2013 No site visit


BBA Inc.	Open Pit mine design, processing plant design, site infrastructure, capital costs, operating costs, financial analysis and overall integration.	David Runnels Patrice Live Colin Hardie	October 11, 2012 October 11, 2012 June 16, 2011



Golder Associates Ltd.	Underground mine design, capital cost and operating costs.	Donald Tolfree	October 11, 2012

AMEC Environment & Infrastructure	Tailings, waste rock and water management, closure plan. Open Pit and Underground geomechanics.	David Ritchie Sheila Daniel Adam Coulson	October 11, 2012 May 19, 2011 January 24-27, 2012

Other Study contributors included Merit Consultants Inc., who provided the schedule, support for the capital cost estimate and constructability. Wayne Clark, of SanZoe Consulting Inc., provided consulting for applications to IESO (Independent Electricity System Operator) and Hydro One Network, energy costs and technical assistance for the high voltage transmission line. Rob Frenette of TBT Engineering Ltd., provided consulting for the Highway 600 reroute and the East Access Road construction.

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1.3

Key Outcomes

Key outcomes from the Feasibility Study are summarized in Table 1-2, Table 1-3 and Table 1-4, and include the following:

•

Proven and Probable Mineral Reserves for combined open pit and underground operations total 116.3 Mt grading 1.08 g/t Au and 2.76 g/t Ag. The open pit mine reserves represent 113.3 Mt (0.97 g/t Au and 2.65 g/t Ag) and the underground mine represents 3.1 Mt (5.07 g/t Au and 6.69 g/t Ag);

•

Planned production rate of 21,000 tpd processing throughput, and a mine life of 16 years (not including two (2) years of pre-production) or 18 years including pre-production years;

•

Low-grade ore stockpiles created during the first nine (9) years will be processed in Years 11 to 16;

•

Planned overburden removal of 80M tonnes over a period of seven (7) years;

•

The open pit operating stripping ratio over the life-of-mine (“LOM”) is 2.81 (waste to reserves ratio) and excludes overburden and capitalized waste;

•

The open pit stripping ratio (LOM) is 3.10 (waste to reserves ratio) and excludes overburden material; and

•

The overall open pit stripping ratio (LOM) is 3.80 (waste and overburden to reserves ratio).

Table 1-2: Key Outcomes – Gold Production


Parameter	Units	Outcome
Year 1 to 5 Average Gold Grade	g/t		1.50
Year 1 to 10 Average Gold Grade	g/t		1.46
Average Gold Grade (LOM)	g/t		1.08
Average Open Pit Gold Grade (LOM)	g/t		0.97
Average Underground Gold Grade (LOM)	g/t		5.07
Total Gold Production from Open Pit (LOM)	koz.		3,188
Total Gold Production from Underground (LOM)	koz.		457
Total Gold Production (LOM)	koz.		3,645
Year 1 to 5 Average Annual Gold Production¹	koz.		330
Year 1 to 10 Average Annual Gold Production¹	koz.		326
Average Process Plant Gold Recovery (LOM)	%		90.4

¹	Production after mining and processing.

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Table 1-3: Key Outcomes – Silver Production


Parameter	Units	Outcome
Year 1 to 5 Average Silver Grade	g/t		2.87
Year 1 to 10 Average Silver Grade	g/t		3.19
Average Silver Grade (LOM)	g/t		2.76
Average Open Pit Silver Grade (LOM)	g/t		2.65
Average Underground Silver Grade (LOM)	g/t		6.69
Total Silver Production from Open Pit (LOM)	koz.		6,186
Total Silver Production from Underground (LOM)	koz.		430
Total Silver Production (LOM)	koz.		6,615
Year 1 to 5 Average Annual Silver Production¹	koz.		441
Year 1 to 10 Average Annual Silver Production¹	koz.		494
Average Process Plant Silver Recovery (LOM)	%		64.1

¹	Production after mining and processing.

Table 1-4: Key Outcomes – Capital Expenses and Financial Results¹


Parameter	Units	Outcome
Initial 5-Year Gold Cash Cost²	USD/oz.		413
Initial 10-Year Gold Cash Cost²	USD/oz.		468
Gold Cash Cost (LOM)²	USD/oz.		544
Open Pit Initial Project Capital (incl. Process Plant and Infrastructure)	$M		713.3
Open Pit-Sustaining Capital³	$M		321.9
Underground – Development Phase Capital³	$M		67.8
Underground – Sustaining Capital³	$M		94.6
Total Royalty Payments	$M		63.3
After-tax Net Present Value @ 5% Discount Rate	$M		931
After-tax Internal Rate of Return	%		23.7
Simple After-tax Payback Period	Years		3.2

The financial analysis was conducted using long term consensus averages of USD $1,400/oz. Au and USD $25/oz. Ag. The United States to Canadian dollar exchange rate was assumed to be USD $1:00:CAD $1.00 during the first two (2) pre-production years and USD $1:00:CAD $1.07 during operation. A discount rate of 5% was used. Cash costs are calculated by recording the mining cost of stockpiled material in the periods in which the material is processed and revenue recognized, in accordance with the IFRS.

²	Includes silver credits and royalty payments. Operating costs for waste quantities above the project’s average stripping ratio of 3.10 have been capitalized.

³	Funded by internal cash flows.

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1.4

Property Description

The Rainy River Gold Project is situated in the southern half of Richardson Township, approximately 50 km northwest of Fort Frances in northwestern Ontario, Canada. The Rainy River Gold Project property is approximately 162 km south of Kenora and 418 km west of Thunder Bay.

The Rainy River Gold Project property comprises a portfolio of 151 patented mining rights and surface rights land claims, including three (3) leasehold interest mining rights land claims and 81 unpatented mining claims covering an aggregate area of 16,697 hectares in the townships of Fleming, Mather, Menary, Potts, Richardson, Senn, Sifton, and Tait. All claims are in good standing and have sufficient work assessment credits available to maintain this good standing for several years.

The Rainy River Gold Project property is divided into two (2) main physiographical regions separated by a distinct northwest to southeast divide, locally termed the Rainy Lake - Lake of the Woods Moraine. The bedrock is overlain by glacial till, which is in turn overlain by thick silts and clays. Some poorly drained areas are also covered by a thick peat layer that impedes exploration activities. The Rainy River Project area is sparsely populated.

1.5

Accessibility, Climate, Local Resources, Infrastructure and Physiography

The climate is typically continental, with extremes in temperature ranging from plus 35°C to minus 40°C. Average rainfall in the region is approximately 60 cm, while approximately 350 cm of snowfall is recorded annually.

The Project site is easily accessible by a network of secondary all-weather roads that branch off the well-maintained Trans-Canada Highways 11 and 71. Access roads are serviced, allowing year-round access. The Canadian National Railway is located 21 km to the south and runs east-west, immediately north of the Minnesota border. The nearby towns and villages of Fort Frances, Emo and Rainy River are located along this railway line.

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1.6

Project History

The Rainy River Gold Project has attracted exploration interest since 1967. Various companies, including Noranda, International Nickel Corporation of Canada, Hudson’s Bay Exploration and Development and Mingold Resources, operated in the area centred on the Project between 1967 and 1989.

The Ontario Geological Survey undertook geological mapping in 1971, and again from 1987 to 1988, in conjunction with a rotasonic overburden drilling program. Nuinsco Resources Limited (“Nuinsco”) undertook exploration activities between 1990 and 2004, with Rainy River continuing from 2005 onwards.

Nuinsco drilled a series of widely spaced reverse circulation drill holes from 1994 to 1998, defining a 15 km long “gold-grains-in-till” dispersal train emanating from a thickly overburden-covered, 6 km² “gold-in-bedrock” anomaly. Nuinsco completed a series of diamond drilling programs to assess the mineral potential of the above anomalies, which led to the initial discovery of the 17 Zone in 1994. Nuinsco subsequently discovered the 34 Zone in 1995 and the 433 Zone in 1997. Between 1994 and 1998, Nuinsco drilled 597 reverse circulation holes and 217 diamond drill holes (49,515 m); these were mostly in the Richardson area. The 34 Zone was further drill-tested between 1999 and 2004.

In June 2005, Rainy River completed the acquisition of a 100% interest in the Project from Nuinsco. That same year, Rainy River relogged key sections of the historical core drilled on the Project property and then input all of the data into a GIS database. Rainy River subsequently drilled in excess of 100 reverse circulation holes in three (3) phases to better define the “gold-in-till” and “gold-in-bedrock” anomalies.

Between 2005 and 2007, 209 diamond drill holes, totalling 95,340 m, were drilled. In April 2008, a mineral resource estimate was completed by Caracle Creek International Consulting. In 2009, SRK prepared a mineral resource statement incorporating information from an additional 112 core boreholes (59,719 m), drilled during 2008. In early 2010, SRK Consulting (Canada) Inc. (“SRK”) prepared a revised mineral resource statement to incorporate information from 124 core boreholes (68,453 m), drilled on the Project during 2009.

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On February 24, 2011, SRK updated the mineral resource statement to incorporate information from 163 core boreholes (84,648 m), drilled on the Project during 2010. A subsequent update, dated June 29, 2011, was prepared by SRK to consider additional drilling (50 boreholes, 26,509 m) completed on the Project between December 2010 and February 27, 2011. This resource update was the basis for the Preliminary Economic Assessment (“PEA”) announced in the November 9, 2011 press release and published on SEDAR on December 23, 2011.

A PEA Update was announced in the August 29, 2012 press release and published on SEDAR on October 12, 2012.

This Feasibility Report is based on SRK’s most recent Mineral Resource Statement disclosed publicly by Rainy River on October 10, 2012. The gold mineral resources are based on drilling data received up to June 10, 2012 and include 1,435 drill holes (662,849 m).

1.7

Geological Setting and Mineralization

The Rainy River Gold Project falls within the 2.7 billion year old Rainy River Greenstone Belt that forms part of the Wabigoon Subprovince. The Wabigoon Subprovince is a 900 km long east-west trending area of komaiitic to calc-alkaline metavolcanics that are in turn succeeded by clastics and chemical sediments. Granitoid batholiths have intruded into these rocks, forming synformal structures in the supracrustals that often have shear zones along their axial planes.

The Wabigoon basement rocks and remnant Mesozoic cover sediments are overlain by Labradorian till of northeastern provenance. This till has been found to contain anomalous concentrations of gold grains, auriferous pyrite and copper-zinc sulphides. It is overlain by a glaciolacustrine clay and silt horizon and by argillaceous and calcareous Keewatin till of western provenance.

The Rainy River Gold Project is primarily underlain by a series of tholeiitic mafic rocks that are structurally overlain by calc-alkalic intermediate to felsic metavolcanic rocks. Intermediate dacitic rocks host most of the gold mineralization. At a regional scale, the strongest and earliest deformation event produced a well-defined penetrative fabric. This foliation is approximately

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parallel to the trend of the metavolcanic rocks that strike at approximately 120 degrees and dip 50 to 70 degrees to the south.

Structural geology studies by SRK suggest that the current geometry and plunge of the gold mineralization is the result of high strain deforming features associated with gold mineralization and rotating the mineralization plunge parallel to the stretching direction.

1.8

Deposit Types

Four (4) main styles of mineralization have been identified on the Rainy River Gold Project:

•

Moderately to strongly deformed, auriferous sulphide and quartz-sulphide stringers and veins in felsic quartz-phyric rocks (e.g., ODM/17, Beaver Pond, 433 and HS Zones);

•

Deformed quartz-ankerite-pyrite shear veins in mafic volcanic rocks (e.g., CAP/South Zone);

•

Deformed sulphide-bearing quartz veinlets in dacitic tuffs/breccias hosting enriched silver grades; and

•

Copper-nickel-platinum group metals mineralization hosted in a younger mafic-ultramafic intrusion (34 Zone).

The gold mineralization is interpreted as a hybrid deposit type, consisting of an early gold-rich volcanogenic sulphide mineralization overprinted by shear-hosted mesothermal gold mineralization. Magmatic-hydrothermal mineralization occurs within the main auriferous zones and crosscuts the volcanogenic sulphide mineralization and the later mesothermal gold mineralization associated with the regional deformation.

1.9

Exploration

In August 2012, exploration drilling discovered a significant new gold and silver zone approximately 1 km east of the proposed Open Pit boundary of the Rainy River Gold Project. Drill hole NR121258 intersected 2.2 g/t gold and 38.5 g/t silver over 18.5 m, including 6.0 g/t gold and 83.9 g/t silver over 3.0 m at a vertical depth of 210 m. This new zone, named the Intrepid Zone, clearly demonstrates the potential for new discoveries along strike of known mineralization. The new zone was discovered after Mobile Metal Ion (“MMI”) geochemistry identified areas of anomalous gold over the prominent magnetic low trend that hosts the majority of the Rainy River Gold Project

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deposits. The new zone contains disseminated and fracture-related mineralization including 2-3% pyrite and variable amounts of sphalerite, galena and chalcopyrite. Both electrum and visible gold have also been identified in core. A total of 102 holes have now penetrated the Intrepid Zone over a 410 m strike length, and have traced the mineralization down dip for 450 m. Four diamond drill rigs will continue to focus on expanding the thickest portions of Intrepid in the plunge direction, and on better defining the zone. National Instrument 43-101 compliant resource estimate is expected in the third quarter (“Q3”) of 2013. The zone provides a target for further drill delineation and a potential future mining opportunity for Rainy River.

The newly discovered Intrepid Zone has not been included in any calculations for mining, plant recoveries or financials and does not affect the content of this Report.

1.10

Drilling

Nuinsco and Rainy River drilled a combined total of 1,435 core boreholes (662,849 m) between 1994 and 2012. Prior to 1999, Nuinsco also drilled several reverse circulation boreholes to sample basal till and bedrock for exploration targeting. Reverse circulation drilling data was not used for resource estimation.

Rainy River drill programs are designed and conducted by an experienced exploration team under the supervision of a Project Manager and the Vice President, Exploration. The drill procedures used by Nuinsco are not well documented. SRK cannot comment on the procedures used by Nuinsco.

SRK is of the opinion that the drilling procedures adopted by Rainy River are consistent with industry best practices and the resulting drilling pattern is sufficiently dense to interpret the geometry and boundaries of the gold mineralization with confidence.

1.11

Sampling Method, Approach and Analyses

There are no records describing the sampling method and approach used by Nuinsco during their 1994 to 2004 drilling program. Previously unsampled intervals within mineralized parts of the

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Nuinsco drill core have been identified and selectively sampled by Rainy River. This additional sampling was incorporated into the drill database.

Rainy River used industry best practices to collect, handle and assay core samples collected during the 2005 to 2012 period. All drilling and sampling was conducted by appropriately qualified personnel under the direct supervision of appropriately qualified geologists.

From early 2005 to late 2006, Rainy River used the accredited ALS Minerals Laboratories (“ALS”) in North Vancouver, British Columbia for sample preparation and analyses. From late 2006 to January 2011, samples were sent primarily to the accredited Accurassay Laboratory (“Accurassay”) facility in Thunder Bay, Ontario. Accurassay used industry standard preparation procedures and standard fire assay procedures for precious metals analyses and aqua regia digestion and atomic absorption spectrometry for metal analyses. In February 2011, Rainy River began using ALS as the primary laboratory for the Project. ALS is accredited by the Standards Council of Canada and its quality management system is accredited to ISO 9001:2008.

Rainy River has partly relied on the accredited laboratory’s internal quality control measures; however, they have also implemented external analytical quality control measures, consisting of inserting control samples (blanks and certified reference material, and field duplicates) with each batch of core drilling samples submitted for assaying. In the opinion of SRK, the field sampling and assaying procedures used by Rainy River meet industry best practices.

1.12

Data Verification

In accordance with National Instrument 43-101 guidelines, Glen Cole, P.Geo., visited the Project property on numerous occasions since 2008. Prior to the Feasibility Study press release, his last visit was from March 24 to 26, 2010. SRK was given full access to all relevant Project data.

SRK conducted a series of routine verifications to ensure the reliability of the electronic data provided by Rainy River. SRK ensured the validation of all tables using Gemcom validation tools that check for gaps, overlaps and out of sequence intervals. SRK summarized the assay results for the external quality control samples for the December 2011 to June 2012 drilling program on time

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series, plots bias charts, quantile-quantile and relative precision plots. Subsequent to the Feasibility Study press release, Glen Cole visited the site from April 30 to May 2, 2013.

Since Rainy River commissioned ALS in February 2011, there has been an overall improvement in the performance of quality control samples. It is SRK’s opinion that gold grades can be reasonably reproduced, suggesting that the assay results reported by the primary assay laboratories are generally sufficiently reliable for the resource estimation used in this Feasibility Study.

1.13

Metallurgical Testwork

1.13.1

Historical Testwork

Initial metallurgical testwork was carried out at SGS Canada Inc. (“SGS”) in Lakefield, Ontario from 2008 to 2011 and was the basis for the PEA Update publicly issued on October 12, 2012. The testwork included mineralization, comminution, gravity separation, flotation, flotation concentrate leaching and whole rock leaching. Results from the testwork indicated that the material was moderately hard. The recovery for a flotation concentrate circuit was estimated at 88.5% with the flotation feed ground to 150 µm and the flotation concentrate ground to 15 µm. The recovery for the whole rock leach circuit was estimated at 91.0% when ground to 60 µm.

1.13.2

Mineralogy

Gold deportment studies were performed at SGS on samples from each zone from 2011 and 2012, including: 5 ODM, 2 Z-433, 1 CAP, 1 HS and 1 NZ sample.

•

The samples were composed mainly of non-opaque minerals, with minor amounts of pyrite present;

•

The gold occurs mainly as native gold, electrum and kustelite. Small amounts of Petzite (Ag₃AuTe₂) were also noted;

•

The gold occurs as liberated, attached and locked particles in most of the samples at a grind size of 150 µm except for the CAP sample;

•

The gold grain size was relatively fine in all samples, with coarse gold (>100 µm) noted only in two (2) of the composites (HS and one of the Z-433 samples); and

•

Trace amounts of pyrrhotite were noted in approximately half the samples.

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1.13.3

Flowsheet Determination Testwork

The majority of flowsheet testwork was conducted at SGS. Various tests, such as grindability, were conducted by suppliers.

Gravity Tailings Leaching

Leaching tests were performed on gravity tailings to compare with a flotation concentrate leach option. The tests were done at grind sizes ranging from 50 to 120 µm. The results from the tests showed gold recoveries of approximately 92% could be achieved. Low reagent consumptions during the tests were also noted.

Flotation Concentrate Leaching

The use of a flotation circuit prior to leaching was investigated. In this processing option, the concentrate from the flotation circuit is leached while the flotation tailings are discarded. The results indicated that an overall circuit gold recovery of approximately 87.5% could be achieved by grinding the flotation concentrate to 15 µm. This ultrafine grind had a high energy requirement and the flotation concentrate leach had high reagent consumptions.

Flowsheet Selection

The gravity tailings leach option was selected in favour of the flotation concentrate option based on the testwork results and is therefore the basis for this Feasibility Study. The main reason for this selection was the significant amount of energy associated with regrinding the flotation concentrate and the high cyanide consumption in flotation concentrate leaching, as well as the additional risk associated with ultrafine grinding of this material. All testwork following this decision was based on the cyanide leaching of gravity tailings.

1.13.4

Comminution Tests

A large scale level comminution testwork program was undertaken to determine the sizing of the semi-autogenuous (“SAG”) mill and ball mill. The tests included 21 crushing work index tests (seven (7) at three (3) separate testing facilities), 16 Bond Ball Mill Work Index (“BWi”), 160 ModBond Ball Mill Work Index, 13 JK Drop Weight tests and 175 SAG Mill Comminution tests

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(“SMC”). In addition, seven (7) samples were sent to Starkey and Associates for SAGDesign testing.

Results from the testwork indicated that the Rainy River material is considered to be hard and relatively resistant to breakage. Based on the testwork results, the following design parameters were determined for the Feasibility Study: Crusher Work Index of 25.0 kWh/t; A x b of 24.2; Bond Ball Work Index (200 mesh open side setting) of 15.0 kWh/t; Bond Abrasion Index of 0.25 g. All design parameters were based on the 80^th percentile of testwork results.

The large testwork campaign provides confidence that the values obtained are representative of the deposit.

Grinding Circuit Design

Five (5) methods were used to size the grinding circuit based on the testwork results: Morrell’s Equation, JK SimMet with Bond Equation, JK SimMet with Phantom Cyclone, SAGDesign and Oreway Mineral Consultants (“OMC”) as an outside consultant.

All methods yielded similarl grinding circuit energy requirements. The lowest overall circuit energy requirement was estimated using SAGDesign at 25.45 kWh/t, while Morrell’s Equation yielded the highest circuit energy requirement at 28.64 kWh/t. The combined JK SimMet and Bond Equation method is the recommended design, resulting in a 26.63 kWh/t circuit energy requirement, including 13.2 kWh/t for the SAG mill and 13.0 kWh/t for the ball mill. This matches well with the energy requirement calculated by OMC.

1.13.5

Gravity Separation

Two (2) Gravity Recoverable Gold (“GRG”) tests were performed on Zone composites representing the ODM and Z-433 Zones. The GRG numbers from these tests were 51.2 and 59.3, respectively. This indicated that the addition of a gravity circuit is beneficial to the process.

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1.13.6

Cyanide Leaching

Additional Gravity Tailings Leaching

Additional gravity tailings leaching tests were performed to determine the retention time and grind size for the variability testwork campaign. Based on these results, a grind size of 75 µm and a retention time of 36 h (with sub-sampling at 30 h) were selected for the variability test program.

Additional testwork was performed to verify the effect of cyanide concentration, pre-conditioning, use of oxygen versus air and lead nitrate. The variability leaching testwork was developed based on results from these tests.

Variability Tests

Cyanide leaching was performed on 208 samples (along with 37 repeats) and the results were used to develop grade-recovery curves for both gold and silver. The average residue value for all the samples was 0.10 g/t Au for the zones other than CAP and 0.16 g/t for the CAP Zone. These residues corresponded to a recalculated gold recovery of approximately 90-91% and 66-67% silver recovery.

The following gold residue (“Au res”) and silver residue (“Ag res”) versus head grade equations (exponential based) were developed using the variability testwork results:


Non-CAP Zones		CAP Zone
Au res = 0.0937 • x_Au^0.4223		Au res = 0.2497 • x_Au^1.015
Ag res = 0.01 • x_Au² + 0.29 x_Ag		Ag res = 0.036 • x_Au² + 0.244 x_Ag

The large leaching testwork campaign has provided significant information regarding the metallurgical response of the material throughout the deposit. This level of testwork reduces the risk that design production levels will not be achieved in operation.

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1.13.7

Cyanide Destruction Testwork

The SO₂/air cyanide destruction process was investigated on two (2) composites: initial pit and remaining life-of-mine. The results showed that this process is effective at lowering the weak acid dissociable cyanide (“CN_WAD”) levels to well below 5 ppm. The average reagent consumptions were 4.7, 3.5 and 0.1 g/gCN_WAD for SO₂, lime and copper, respectively.

1.13.8

Environmental Testwork

AMEC Environment & Infrastructure is conducting environmental geochemical characterizations of selected samples, representative of the mine rock and overburden in the vicinity of the proposed Rainy River Gold Project pit, and tailings produced in metallurgical testwork. To date, testing has been carried out on three (3) simulated tailings materials, and a total of 659 deposit-wide mine rock samples, of which 366 represent in-pit non-ore mine rock.

Geochemical studies to-date on mine rock indicates that approximately half the samples may have the potential to produce acid rock drainage. A block model was developed to refine the estimated tonnage of potentially acid generating rock. Generally, metal contents in waste materials are typical for their rock types and the risk for metal leaching under neutral conditions appears to be low. Humidity cell analysis is ongoing on the mine rock and tailings to evaluate the long-term metal leaching characteristics of these materials.

The results of the mine rock and tailings analyses indicate a risk for acid rock drainage from a portion of the Rainy River Gold Project mineral waste in the future, if not appropriately managed. The Rainy River Gold Project design has taken this into account in the operation and closure of the East stockpile and Tailings Management Area (“TMA”).

1.13.9

Testwork Interpretation

Results from the SGS testwork are the basis for the mineral reserve estimate and Feasibility Study. Based on a trade-off study, it was determined that the whole rock leaching option with gravity separation was the most economical alternative compared to the flotation option, and was therefore used as the basis for the Feasibility Study. The main reason for this selection was the

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significant amount of energy associated with regrinding the flotation concentrate and the high cyanide consumption in the flotation concentrate leaching, in addition to risk associated with ultrafine grinding of this material. All subsequent testwork was based on cyanide leaching of the gravity tailings.

The extensive grinding testwork campaign has allowed for definition of the overall hardness of each zone and indicated that there are several portions of the deposit that will have high energy requirements and this will be reflected in the design of the process plant. The strong correlation between the four (4) methods used to size the grinding circuit provides a good level of confidence in the sizing of the SAG and ball mill.

The process is expected to yield an overall gold recovery of approximately 90-91% and a silver recovery of approximately 66-67% over the life-of-mine without considering solution losses. When considering solution losses, the gold recovery decreases slightly by 0.3% while the silver recovery drops to approximately 64%. The grind size chosen for this study was 75 µm, based on a cost versus revenue study performed by BBA. Eight (8) agitated leach tanks will be used in a series arrangement, allowing for a 30 hour retention time. Seven (7) carbon-in-pulp (“CIP”) tanks will be required with a total circuit retention time of 112 minutes. The gold recovery varies significantly throughout the deposit and the CAP Zone has considerably lower gold recoveries than the other zones. As such, the mine plan is based on stockpiling most of the CAP Zone material and processing it at the end of the mine life.

1.14

Mineral Resource Estimate

The mineral resource model presented herein represents the eighth resource evaluation for the Rainy River Gold Project. The resource estimate was completed by Dorota El-Rassi, P.Eng. (APEO #100012348) and Glen Cole, P.Geo. (APGO #1416) from SRK Consulting, both independent Qualified Persons for the purpose of National Instrument 43-101. The effective date of the Mineral Resource Statement is October 10, 2012.

The Rainy River database comprises data from 1,435 core boreholes (663 km), drilled by Nuinsco and Rainy River. All exploration information is located using a UTM grid (NAD 83 datum, 15 Zone). Resource modelling and grade estimation were conducted in this UTM coordinate space.

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SRK updated a series of previously constructed 3D wireframes used to constrain the extent of the gold mineralization, considering structural features, lithology, alteration, geochemical indices, as well as grade trends. From these wireframes, 13 distinct mineralized zones with further domains within these were constructed and used as hard boundaries for estimating the mineral resources.

For geostatistical analysis, variography and grade estimation, raw assay data were composited to 1.5 m lengths. The impact of capping was analyzed and capping levels were adjusted for each resource domain and each metal separately. Capping was applied to the composited data. An unrotated block model was created to cover the entire extent of the Rainy River deposit area. Block size was set at 5 x 5 x 5 m.

Variography was undertaken to characterize the spatial continuity of gold and silver within each resource domain, and to assist with the selection of estimation parameters. Metal grades were estimated using ordinary kriging as the principal estimator, separately, in each domain from capped composite data within that domain. Grades in domains 601 to 605 were estimated using an inverse distance algorithm. Three (3) estimation passes using search neighborhoods sized from variography results were used to populate the block models. The first estimation pass generally considered search neighborhoods adjusted to half or full variogram ranges, with the search ellipse then doubled for the second and third estimation pass.

The mineral resources were classified as Measured, Indicated and Inferred, primarily determined on the basis of block distance from the nearest informing composites and on variography results. The classification strategy was based on gold data alone and also considered the geological setting of the Project.

SRK considers that portions of the Rainy River Gold Project mineralization are amenable for open pit extraction, while other parts of the deposits could be extracted using an underground mining method. To assist with determining which portions of the modelled mineralization show “reasonable prospect for economic extraction” from an open pit, and to assist with selecting reasonable reporting assumptions, SRK used a pit optimizer to develop conceptual open pit shells. The block model was also reviewed to determine which portions of the gold mineralization have “reasonable prospects for economic extraction” from an underground mine. The mineral resource

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for the Rainy River Gold Project is reported at two (2) cut-off grades (Table 1-5). The effective date of the Mineral Resource Statement is October 10, 2012. Open pit mineral resources are reported at a cut-off grade of 0.35 g/t gold, whereas underground mineral resources are reported at a cut-off grade of 2.5 g/t gold.

Table 1-5: Mineral Resource Statement¹, Rainy River Gold Project, Ontario,

SRK Consulting (Canada) Inc., October 10, 2012


	Quantity		Grade				Metal
	Quantity		Au		Ag		Au		Ag
Category	‘000 t		g/t		g/t		‘000 oz.		‘000 oz.
In Pit Mineral Resources²(cut-off grade: 0.35 g/t gold)
Measured		27,550		1.32		1.90		1,168		1,681
Indicated		112,271		1.11		2.51		4,012		9,048
Measured and Indicated		139,821		1.15		2.39		5,180		10,728
Inferred		19,353		0.88		1.40		550		870
Out of Pit Mineral Resources²
Indicated		14,466		0.80		3.84		373		1,785
Inferred		73,555		0.68		2.53		1,610		5,980
Underground Mineral Resources²(cut-off grade: 2.5 g/t gold)
Measured		88		4.97		2.76		14		8
Indicated		4,148		4.50		6.12		600		816
Measured and Indicated		4,236		4.50		6.05		614		824
Inferred		897		4.18		4.63		120		134
Combined Mineral Resources: In Pit, Out of Pit and Underground²
Measured		27,638		1.33		1.90		1,182		1,689
Indicated		130,885		1.18		2.77		4,985		11,649
Measured and Indicated		158,523		1.21		2.62		6,167		13,338
Inferred		93,805		0.75		2.32		2,280		6,984

Mineral resources are reported by relative conceptual pit shells. On average, the open pit extends to an elevation of 500 m below surface. Mineral resources are not mineral reserves and do not have a demonstrated economic viability. All figures are rounded to reflect the relative accuracy of the estimate. Figures may not add due to rounding. All assays have been capped where appropriate. Qualified persons - The mineral resource statement was prepared by Dorota El-Rassi, P.Eng. (APEO #100012348) and Glen Cole, P.Geo. (APGO #1416), of SRK, both “independent qualified persons” as that term is defined in National Instrument 43-101. Rainy River’s exploration program in Richardson Township is being supervised by Kerry Sparkes, P.Geo. (APEGBC #25261), Vice-President, Exploration and a Qualified Person as defined by National Instrument 43-101. The Company continues to implement a rigorous QA/QC program to ensure best practices in sampling and analysis of drill core. The estimates of mineral resources may be materially affected by environmental, permitting, legal, title, taxation, sociopolitical, marketing, and other relevant issues.

Open pit mineral resources are reported at a cut-off grade of 0.35 g/t gold, underground mineral resources are reported at a cut-off grade of 2.5 g/t gold based on a gold price of USD $1,100 per ounce, a silver price of USD $22.50 per ounce and a foreign exchange rate of CAD $1.10 to USD $1.00. Metallurgical recoveries include 88% for gold in the open pit resources and 90% for gold in the underground resources with a silver recovery of 75%.

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1.15

Open Pit and Underground Mine Design

The evaluation of the open pit reserves was supervised by Patrice Live (BBA Inc.) and the underground reserves were evaluated by Donald Tolfree (Golder Associates Ltd.). The combined open pit and underground reserves were generated from the block model provided by SRK in October, 2012. The block model consists of small blocks measuring 5 m E x 5 m N and 5 m in elevation.

The Feasibility Study assumes both open pit and underground mining methods will be used for resource extraction. The mining methods and production capacity have been chosen to match a milling throughput rate of 21,000 tpd (20,000 tpd from the open pit and 1,000 tpd from underground when full production is achieved). Both the open pit and underground operations will deliver material to a common gyratory crusher for primary size reduction and delivery to the processing plant. Utilization of Rainy River’s mining equipment and personnel is envisioned for the development of the open pit as well as for the removal of overburden. Specialized underground contractors will be used for the development of the main underground ramp and vertical excavations. Rainy River equipment and personnel will be used to complete the underground development and for all production activities.

1.15.1

Open Pit Mine

Open Pit Mine Design

Surface mining of the Rainy River Gold Project will follow the standard practice of an open pit operation, with conventional drill and blast, load and haul cycle using a drill/truck/shovel mining fleet.

In order to develop an optimal engineered pit design for the Rainy River deposit, an optimized pit shell was first prepared using the Lerchs-Grossman 3D (“LG 3D”). routine in MineSight. With defined pit optimization parameters including gold and silver prices, mining, processing and other indirect costs, Au and Ag recoveries for each ore type (as determined from metallurgical testwork), pit slopes (by AMEC based on geotechnical pit slope study) and imposed constraints, the pit optimizer searches for the pit shell with the highest undiscounted cash flow. The main pit optimization parameters used in the LG 3D routine are listed in Table 1-6.

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Table 1-6: Pit Optimization Parameters


Type of Activity	Unit	Values
Mining Cost	$/t mined	1.89
Processing Cost	$/t milled	8.73
General and Administration Cost	$/t milled	1.00
Refining Cost	$/t milled	1.50
Au Recovery	%	As per equation
Ag Recovery	LOM average %	64.1
Gold Selling Price	USD/oz.	Varied from 200 to 1250
Silver Selling Price	USD/oz.	25
Exchange Rate	CAD/USD	1.05
Overall Pit Slope Angle	degree	Varied from 38 to 51
Overall Overburden Slope Angle	degree	16
Surface Limitation from Pinewood Creek	M	100
Surface Limitation from Bayfield’s Properties	M	0
Depth/elevation constraint	MASL	–50

Using the technical and economic parameters described previously, the MineSight LG 3D pit optimizer tool was run to produce a series of pit shells at different gold prices. Once the series of pit shells were generated, total material moved, total open pit resource, NPV and stripping ratios were evaluated to identify the optimum pit shell. The selected optimized pit for this Study is based on a gold price of USD $800/oz. with the highest undiscounted cash flow and a mine life of approximately 15 years.

The optimum pit shell selected was used as a guide to carry out the detailed mine design using the design parameters, shown in Table 1-7. Access to the pit uses a 10% gradient ramp, 33 m wide, to accommodate large size trucks for double-traffic lane haulage. Overburden and rock slope configurations and angles were all based on geomechanical design recommendations provided by AMEC.

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Table 1-7: Open Pit Mine Design Parameters


Optimization Parameter	Unit	Value
Benching Arrangement Height	m	3 x 10
Ramp Road Width – Two (2) lanes	m	33
Ramp Road Width – Single (2) lane	m	20
Ramp Grade in Rock	%	10
Overall Pit Slope Angle	degree	Varied from 40 to 56.3
Overall Overburden Slope Angle		Varied from 3.25H:1V to 4H:1V

Using the parameters described in Table 1-7 and a gold price of USD $1,250/oz., an equivalent (Au eq) open pit cut-off grade (“COG”) of 0.30 g/t Au eq was calculated. Dilution and material loss were estimated using digitized mining polygons representing the mining area. The open pit dilution and mine recovery rates were estimated to be 9.7% at 0.22 g/t Au and 1.31 g/t Ag and 95%, respectively. The final pit contains 113.2 Mt at 0.97 g/t Au and 2.65 g/t Ag of measured and indicated reserves. The average life-of-mine design overall strip ratio (including overburden and waste rock) is 3.8:1, with an average life-of-mine open pit strip ratio (waste rock only) of 3.1:1. In order to access the open pit, approximately 80 million tonnes of overburden must be removed. Excluding capitalized waste, identified as operating costs for waste quantities with a stripping ratio above 3.1:1, the project’s operating strip ratio is 2.8:1.

1.15.2

Underground Mine

Underground Mine Design

To optimize the value of the underground mine, multiple underground designs at various COGs and layouts were investigated. The final design is based on a COG of 3.5 g/t Au eq and includes 3.1 million tonnes of ore grading 5.07 g/t Au and 6.69 g/t Ag. These reserves will be mined with a combination of LHOS and mechanized CAF mining methods employing both cemented rockfill (“CRF”) and unconsolidated rockfill (“RF”). The total waste rock extracted from the underground mine is approximately 2 million tonnes and a portion of this will be used for rockfill.

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CAE’s Mineable Shape Optimizer (“MSO”) (a plugin to Datamine) was used to estimate the potential in-situ longhole tonnage. The resource model was evaluated with a seed shape that is 5 m wide, 5 m long, 25 m high (this measurement is vertical, not dipped) and has a hangingwall (“HW”) dip of 60 degrees, to seek stope shapes that had an average grade above the COG. The resulting stope shapes that were clustered together were combined into stoping areas and smoothed to provide a more realistic hangingwall and footwall (“FW”) profile, and grade estimate. Stope shapes that were not clustered together or clusters that did not contain enough ounces to pay for access infrastructure, were considered orphans and removed from the reserve. Approximately 80% of the underground reserves are in the LHOS stopes.

Shallower dipping areas that were above the COG were evaluated with the CAF mining method. The CAF stopes were designed using the resource model blocks (5 m wide by 5 m high) and simple economics were used to determine if an area should be added to an existing CAF lift. If a mining area was mineable by both CAF and LH mining methods, the method that either produced the highest margin per tonne ore or matched the surrounding method was chosen. Approximately 20% of the underground reserves are in the CAF stopes.

Hangingwall dilution of the LH stoping areas was estimated by extruding the stoping area shapes according to the geotechnical recommendations provided by AMEC. Additional LH stope dilution is from the CRF mucking floor and vertical CRF walls (secondary stopes only). Overall, it is estimated that the LH stopes will have approximately 10% dilution grading 1.56 g/t Au and 1.28 g/t Ag. Dilution in the CAF areas occurs from the RF mucking floor and side wall sloughing and is estimated to be 9% with a grade of 0.61 g/t Au and 4.16 g/t Ag. Both mining methods assume a 95% mining recovery.

1.15.3

Open Pit and Underground Reserves

The combined open pit and underground mine mineral reserves are summarized in Table 1-8. This summary is reported at two (2) gold equivalent COGs. Open pit mine reserves are reported at a COG of 0.30 g/t Au eq, whereas underground reserves are reported at a COG of 3.5 g/t Au eq. The summary of mineral reserves includes both dilution and mining recovery factors.

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Table 1-8: Open Pit and Underground Proven and Probable Mineral Reserves (April 10, 2013)^1,2,3,4


Resources Category	Tonnage (Mt)		Au Grade (g/t)		Ag Grade (g/t)		Au (In-Situ oz.)		Ag (In-Situ oz.)
Open Pit
Proven		27.7		1.14		1.94		1,014,584		1,727,979
Probable		85.5		0.91		2.88		2,510,641		7,918,793

TOTAL		113.2		0.97		2.65		3,525,225		9,646,772

Underground
Proven
Probable		3.1		5.07		6.69		506,283		668,240

TOTAL		3.1		5.07		6.69		506,283		668,240


Total Combined
Proven		27.7		1.14		1.94		1,014,584		1,727,979
Probable		88.6		1.06		3.01		3,016,924		8,587,034

TOTAL		116.3		1.08		2.76		4,031,508		10,315,013

Open pit reserves have been estimated using a cut-off grade of 0.30 g/t gold-equivalent, and underground reserves have been estimated using a cut-off grade of 3.5 g/t gold-equivalent. Open pit reserves have been estimated using a dilution of 9.7% at 0.22 g/t Au and 1.31 g/t Ag, and underground reserves have been estimated using a CAF dilution of 9% at 0.61 g/t Au and 4.16 g/t Ag and LH dilution of 10% at 1.56 g/t Au and 1.28 g/t Ag. Open pit reserves have been estimated using a mine recovery of 95%, and underground reserves have been estimated using a mine recovery of 95%.

Qualified persons - The mineral reserve statement was prepared by Patrice Live (OIQ #38991) of BBA, and Donald Tolfree (APEGBC #32557), of Golder Associates, both “independent qualified persons” as that term is defined in National Instrument 43-101. Rainy River’s engineering assessment in Richardson Township is being supervised by Garett Macdonald, P.Eng. (PEO #90475344), Vice-President, Operations and a Qualified Person as defined by National Instrument 43-101.

Reserves are derived from the October 10, 2012 Resource Statement, prepared by Dorota El-Rassi, P.Eng. (APEO #100012348) and Glen Cole, P.Geo. (APGO #1416), of SRK, both “independent qualified persons” as that term is defined in National Instrument 43-101. Rainy River’s exploration program in Richardson Township is being supervised by Kerry Sparkes, P.Geo. (APEGBC #25261), Vice-President, Exploration and a Qualified Person as defined by National Instrument 43-101. The Company continues to implement a rigorous QA/QC program to ensure best practices in sampling and analysis of drill core.

⁴	The mineral reserve estimate may be materially affected by environmental, permitting, legal, title, taxation, sociopolitical, marketing, and other relevant issues.

1.16

Mining Methods

The Feasibility Study assumes both open pit and underground mining methods will be used for ore extraction. The mining methods and production capacity have been chosen to match a milling throughput rate of 21,000 tpd.

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Open Pit and Underground Geotechnical Design

The feasibility level site investigation work, open pit slope design criteria for slope stability, underground mine design criteria for stope stability, sequencing, ground support and backfill, were performed by AMEC (2012F; 2012G, 2013A; 2013B; 2013C, 2013D, 2013E, 2013F) supervised by Adam Coulson (AMEC) and provided to BBA for open pit design and to Golder Associates for underground mine design.

During the 2012 AMEC geomechanical drilling campaign, assessment of various rock structural domains was based on the analysis, ten (10) NQ-sized boreholes with geomechanical logging of oriented core and packer testing down the holes to understand the hydrogeological characteristics. These boreholes were oriented in various azimuths, dipping on average at 65° in order to intersect potential open pit walls, know geological features and the three main underground zones of the ODM/17 (West, Central and East) for a combined drilling and logging length of around 4.5 km. This investigation work was supported with acoustic televiewer surveys by DGI Geoscience Inc. of ten (10) exploration boreholes to confirm the orientation of the major joint sets at depth.

A total of 176 core samples were collected for laboratory strength testing by AMEC, with 268 test specimen prepared. From these, 117 specimens were tested for uniaxial compressive strength (“UCS”), 42 for triaxial strength, 100 for Brazilian tensile strength, and nine (9) for direct shear tests of open joints. Elastic properties (Young’s Modulus and Poisson’s ratio) were also assessed on 20 specimens. Additionally, for backfill strength assessment under various binder mixes by AMEC, two (2) bulk samples of potentially acid generating and potentially non acid generating rock in excess of 1,000 kg each were sampled and crushed for short and long-term strength testing.

Based on the field and laboratory investigations, open pit slope stability design criteria for optimised bench face angles, inter-ramp angles and overall slope angles was performed using probabilistic kinematic and limit equilibrium structural analyses, including numerical analyses. Similarly the underground stope dimensions, ground support requirements, backfill strength and effects of stope sequencing on development and ground support levels, was performed with a combination of numerical stress modeling and recognized empirical design tools.

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1.16.1

Open Pit Operations

Conventional open pit mining has been chosen as the primary method to extract the Rainy River deposit because of the deposit’s proximity to the surface. The open pit mine production rate is planned to be approximately 21,000 tpd of mill feed material initially, and then 20,000 tpd when the 1,000 tpd underground operation comes on line. Production is based on a declining COG strategy, whereby low-grade material (0.3 to 0.6 g/t Au) will be stockpiled near the primary crusher for future processing. The objective of this strategy is to obtain a higher average milling grade in the early years to maximize the Project’s cash flow and economics. The low-grade stockpiles will be reclaimed gradually at the end of the mine life or on an as-needed basis.

The primary equipment fleet at the peak of operations consists of: three (3) hydraulic shovels, one (1) wheel loader, 19 haul trucks, and a fleet of support equipment. Production drilling will be carried out by diesel-powered track-mounted units. Operating bench heights of 10 m have been planned for mining operations. Over the life of the mine, a total of 350.6 Mt of waste rock and 80.0 Mt of overburden will be moved. It is anticipated that all overburden will be removed within the first seven (7) years of mine operation. At the peak of open pit mining, a workforce of 299 persons (staff and hourly employees) will be required. Mined waste and overburden will be stored in nearby stockpiles or used in dam construction activities associated with the TMA.

The open pit mine schedule is based on pushback designs by BBA and recommendations presented in Whittle Consulting’s Enterprise Optimization (“EO”) report. The EO process is an integrated approach to maximizing the Net Present Value (“NPV”) of a mining business by simultaneously optimizing 10 different mechanisms across the mining value chain. It was conducted to support the PEA Update and this Feasibility Study.

The open pit mine schedule is shown in Table 1-9. Open pit reserves have been estimated using:

•

A COG of 0.30 g/t equivalent Au;

•

Dilution of 9.7% at 0.22 g/t Au and 1.31 g/t Ag; and

•

Mine recovery of 95%.

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Table 1-9: Open Pit Mine Schedule


Year	Process Cut-off Grade		Mine to Mill						Mine to Stockpile						Stockpile to Mill						Waste		Over- burden		Total Material Moved
Year	(g/t)		(Mt)		Au (g/t)		Ag (g/t)		(Mt)		Au (g/t)		Ag (g/t)		(Mt)		Au (g/t)		Ag (g/t)		(Mt)		(Mt)		(Mt)
2014		0.60		—		—		—		—		0.55		1.74		—		—		—		2.81		8.47		11.28
2015		0.60		—		—		—		0.38		0.61		2.03		—		—		—		11.35		14.21		26.31
2016		0.60		2.87		1.19		3.13		3.47		0.39		2.12		—		—		—		19.67		11.95		41.42
2017		0.60		7.66		1.31		2.99		6.47		0.38		1.94		—		—		—		35.77		6.74		63.12
2018		0.60		7.66		1.35		3.85		5.31		0.36		2.23		—		—		—		31.60		11.82		61.69
2019		0.60		7.52		1.72		2.06		4.12		0.38		1.19		—		—		—		30.25		21.06		67.07
2020		0.50		7.42		1.58		2.36		5.31		0.37		1.78		—		—		—		43.62		5.78		67.44
2021		0.50		7.25		1.36		2.72		5.59		0.36		2.04		—		—		—		47.02		—		65.46
2022		0.45		7.33		1.12		3.38		4.81		0.35		2.15		—		—		—		47.82		—		64.76
2023		0.40		7.31		1.10		5.01		4.05		0.37		2.39		—		—		—		46.13		—		61.54
2024		0.45		7.30		1.15		3.43		3.61		0.33		1.80		—		—		—		28.28		—		42.81
2025		0.45		7.31		1.39		1.88		—		—		—		—		—		—		5.92		—		13.23
2026		0.45		0.49		1.39		1.88		—		—		—		6.82		0.37		1.98		0.40		—		0.89
2027		—		—		—		—		—		—		—		7.37		0.36		2.17		—		—		—
2028		—		—		—		—		—		—		—		7.45		0.36		1.40		—		—		—
2029		—		—		—		—		—		—		—		7.66		0.32		1.92		—		—		—
2030		—		—		—		—		—		—		—		7.66		0.30		2.10		—		—		—
2031		—		—		—		—		—		—		—		6.16		0.52		2.28		—		—		—

TOTAL:		—		70.12		1.34		3.07		43.12		0.37		1.97		43.12		0.37		1.97		350.64		80.03		587.02

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Figure 1-1 shows the final pit design,the initial pit, and resource pit shell.

LOGO

Figure 1-1: Initial Pit, Final Pit and Resource Pit Shell

1.16.2

Underground Operations

At a 3.5 g/t Au eq COG, LHOS and CAF mining methods are proposed. LH mining will mainly follow a traditional transverse primary-secondary pyramid sequence using CRF in the primary stopes and unconsolidated RF in the secondary stopes. There are minor amounts of longitudinal stoping that will use CRF. The LHOS are designed with a 25 m sublevel interval and a hangingwall dip of 60 degrees. The primary stopes are designed to be an average of 10 m wide (across strike) by 10 m long (along strike), and the secondary stopes are designed to be 10 m wide by 20 m long.

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The CAF mining method is proposed in shallower dipping areas that are above the COG. This method involves mining 5 m lifts from the bottom up; the lowest lift of a mining zone is mined first. Once complete, this lift is backfilled with unconsolidated rockfill and then the next 5 m lift is mined on top of the backfill.

The mining areas are accessed by a main ramp that is 6.0 m wide by 5.5 m high with an average grade of 15%. In total, the underground mine will require approximately 27 km of capital excavation, which includes the main ramps, the footwall accesses for the LH stopes and the infrastructure required to support mining activities such as a small underground shop, sumps, and ventilation drifts and raises. Mined waste will either be used as backfill when possible or hauled to the surface waste storage facility. It is estimated that approximately 50% of the underground waste will remain underground as backfill.

The underground mine production rate is estimated to average 1,000 tpd. The pod-like nature and strike length of the mineralized material will allow for multiple working fronts, which will provide increased productivity for the mining methods proposed. A fleet of 5.4 m³ load-haul-dumps (LHDs) and 27 m³ trucks has been selected as the primary underground production equipment. At peak operation, the underground will have approximately 42 pieces of mobile equipment (including production, development and support equipment), will require a workforce of 187 persons (staff, hourly employees and contractors), and have a ventilation capacity of approximately 350 m³/s.

The underground mine schedule is shown in Table 1-10. Underground reserves have been estimated using:

•

COG of 3.5 g/t equivalent Au;

•

Dilution of 9.7% at 1.56 g/t Au and 1.28 g/t Ag for the longhole areas and 9.0% at 0.61 g/t Au and 4.16 g/t Ag for the cut and fill areas; and

•

Mine recovery of 95%.

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Table 1-10: Underground Mine Schedule


	Cut and Fill						Longhole						Development
Year	(kt)		Au (g/t)		Ag (g/t)		(kt)		Au (g/t)		Ag (g/t)		(kt)		Au (g/t)		Ag (g/t)
2014		—		—		—		—		—		—		—		—		—
2015		—		—		—		—		—		—		—		—		—
2016		—		—		—		—		—		—		—		—		—
2017		—		—		—		—		—		—		—		—		—
2018		—		—		—		—		—		—		4		5.33		5.69
2019		—		—		—		142		4.06		4.15		6		5.94		6.81
2020		48		4.03		2.23		201		4.81		4.80		0.7		5.56		2.65
2021		95		3.98		1.86		268		4.86		3.30		48		5.47		2.59
2022		95		4.89		1.90		256		5.99		2.96		19		5.84		3.43
2023		95		4.53		14.26		251		5.89		2.30		14		4.68		2.21
2024		101		4.07		38.49		257		5.98		2.00		3		5.09		1.42
2025		102		4.43		41.70		265		5.66		2.02		—		—		—
2026		91		3.61		30.62		267		5.16		2.06		—		—		—
2027		23		3.63		25.80		269		4.97		3.05		—		—		—
2028		—		—		—		219		4.92		4.90		—		—		—
2029		—		—		—		—		—		—		—		—		—
2030		—		—		—		—		—		—		—		—		—
2031		—		—		—		—		—		—		—		—		—

TOTAL:		6.48		4.22		20.56		2364		5.29		3.04		95		5.44		3.06

Figure 1-2 provides an isometric view of the underground mine and ultimate open pit.

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LOGO

Figure 1-2: Isometric View of the Rainy River Underground Mine

and the Ultimate Pit (not to scale)

1.16.3

Proposed Mine Plan

The proposed Project will be a combined open pit (20,000 tpd) / underground (1,000 tpd) operation with approximately 116 Mt of material being processed over the life-of-mine at a nominal mill daily throughput of 21,000 tpd (7.7 Mtpa). The open pit mine production rate is planned to be approximately 21,000 tpd of mill feed material initially (2016) and then gradually decrease to 20,000 tpd when the 1,000 tpd underground operation comes on line (2018). The mine schedule contains two (2) years of pre-production and envisions a mine operating life of 16 years exclusive of the pre-production period. Development of the underground operation would start in Year 1 (2016) of mine operations and start producing at full rates by Year 6 (2021). Figure 1-3 presents the annual gold and silver production based on the proposed mine plan (open pit and underground) and process plant gold and silver recoveries.

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LOGO

Figure 1-3: Annual Gold and Silver Production (koz.)

Over the life-of-mine, 3,188 koz. of gold and 6,186 koz. of silver will be recovered from the open pit operation (average open pit LOM grade: 0.97 g/t Au and 2.65 g/t Ag) and 457 koz. of gold and 430 koz. of silver will be recovered from the underground operation (average underground LOM grade: 5.07 g/t Au and 6.69 g/t Ag). A total production of 3,645 koz. of gold and 6,615 koz. of silver is expected from the Rainy River Gold Project.

The combined open pit and underground mine schedule is presented in Table 1-11.

Table 1-11: Total Milled OP and UG


	Total Milled UG and OP
Year	(Mt)		Au (g/t)		Ag (g/t)
2014		—		—		—
2015		—		—		—
2016		2.87		1.19		3.13
2017		7.66		1.31		2.99
2018		7.67		1.36		3.85
2019		7.66		1.76		2.11
2020		7.67		1.68		2.43
2021		7.66		1.54		2.73
2022		7.67		1.32		3.35

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	Total Milled UG and OP
Year	(Mt)		Au (g/t)		Ag (g/t)
2023		7.67		1.31		5.03
2024		7.66		1.35		3.85
2025		7.67		1.58		2.41
2026		7.66		0.64		2.31
2027		7.66		0.53		2.27
2028		7.67		0.49		1.5
2029		7.66		0.32		1.92
2030		7.66		0.3		2.1
2031		6.16		0.52		2.28

TOTAL:		116.3		1.08		2.76

1.17

Process Plant

The process plant facility is designed to have an availability of 92% and a capacity of 21,000 tpd (7.7 Mtpa). This includes 1,000 tpd from the underground mine. Average head grade for the plant is 1.08 g/t Au while the design grade is 2.0 g/t Au, in order to accommodate periods of higher grade feed. The average silver head grade is 2.76 g/t Ag while the design grade is 5.0 g/t Ag.

Run-of-mine material will be crushed to 165 mm in a 1,371 mm x 1,905 mm (54” x 75”) gyratory crusher (448 kW) and then stockpiled. The primary crusher building houses the gyratory crusher and the tail-end of the stockpile feed conveyor.

The crushed rock will be withdrawn from beneath the stockpile and ground to approximately 2,400 µm in an 11.0 m x 6.1 m (36’ x 20’), 15,000 kW SAG mill, in closed circuit with a scalping screen and a pebble crusher. The 448 kW (600 HP) pebble crusher will crush the scalping screen oversize to 13 mm. The undersize from the scalping screen will be combined with the ball mill discharge and pumped to a cyclone cluster. The underflow from the cyclone cluster will feed the 7.9 m x 12.3 m (26’ x 40.5’) 15,000 kW ball mill. A portion of the ball mill discharge will be treated by a gravity circuit, which includes an intensive cyanidation unit and dedicated electrowinning cell. The gravity tailings are returned to the grinding circuit while the tailings from intensive cyanidation are sent to the leaching circuit.

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The cyclone overflow will be thickened in a pre-leach thickener. The pre-leach thickener underflow will be pumped to the cyanide leaching circuit with one (1) series of eight (8) 18 m diameter tanks. The leach circuit has been designed for approximately 30 hours retention time. The discharge from the leach circuit will flow by gravity to a carousel-style CIP circuit with seven (7) tanks, where the leached gold will be adsorbed onto the carbon. The slurry containing the loaded carbon will be pumped from the CIP circuit and then screened, and the oversized material (loaded carbon) will be sent to the carbon stripping circuit. The pregnant strip solution will be cooled in a heat exchanger and discharged into an electrowinning cell. The gold and silver will be recovered as a sludge in the electrowinning cells, filtered, dried and then smelted into a doré bar. The coarse spent carbon from the stripping circuit will be reactivated in a kiln, while the fine carbon will be bagged and sold for silver and gold credit.

The tailings from the CIP circuit will be sent to a pre-detox tailings thickener. The underflow from the thickener will be diluted with non-cyanide bearing process water prior to being pumped to cyanide destruction. The SO₂/air process will be used to lower the cyanide level in the tailings to an acceptable level.

A schematic flowsheet of the process is presented in Figure 1-4.

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LOGO

Figure 1-4: Schematic Process Flowsheet

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A steel building will house the grinding area and the gold refinery. Mill reagents, grinding steel and maintenance supplies will be delivered to the site by transport truck and stored in the mill, as required.

In order to minimize fresh water use, process water for the mill facility will be reclaimed from the tailings pond area.

Overall gold recovery of the proposed circuit will be approximately 90.4% for the life-of-mine (LOM) and silver recovery will be approximately 64.1%. It is estimated that 89 people (staff and hourly employees) will be required for the process plant operations.

1.18

Project Infrastructure

A general site layout including the open pit, processing area, tailings management area, emulsion plant, truck shop and various stockpiles is shown in Figure 1-5. Details of the site are described in the following sections.

Site Access

Ontario Provincial Highway 600 currently runs directly through the Project site. The deviation of the road will require existing roads to be upgraded to provincial highway standards and a section of new highway will need to be built.

Site access roads to the TMA and to the explosives plant are pre-existing roads. The road widths will be enlarged to allow space for tailings and reclaim water pipes, as well as light traffic (emulsion tankers and pickup trucks). Existing roads will be resurfaced with crushed stone.

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LOGO

Figure 1-5: General Site Layout

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Buildings

The main administration building will be located at the entrance of the mine site and will house the administration and safety/security staff.

The mine office will be located next to the truck shop and will house the mine, maintenance and engineering office staff. The building will also have dry facilities with lockers.

The plant office will be located on the west side of the process building between the leach tanks and the pre-leach thickener and will be connected to the main building via a short corridor. The building will house the process operations/maintenance office staff, and a dry facility.

The assay laboratory is a separate building and will be equipped for all process plant and mine assaying requirements.

The mine truck shop, for both the open pit and underground operations, will have six (6) maintenance bays, including two (2) bays for auxiliary vehicles and one (1) bay dedicated for welding. The building will include a 1,400 m² warehouse and a mechanical workshop which will also serve as a maintenance area for small vehicles.

The truck wash facility will be located approximately 80 m south of the mine truck shop and will incorporate the tire change bay.

Heating, ventilation and air conditioning (“HVAC”) will be provided for all buildings based on the required working temperatures.

The mine fleet fuel island will be located close to the crusher on the main haul road to the plant and facilities. The tank farm will be located on the service road between the crusher and stockpile. The light vehicle fleet fuel island will be located on the main access road near the warehouse.

Site Utilities

The total power demand of the Project is estimated at 56.9 MW. Electricity will be supplied by a new 17 km long 230 kV power line, to be built and subsequently connected to the existing 230 kV

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Hydro One line connecting Fort Frances and Kenora. The main 230–27.6 kV substation will be located near the concentrator building. The electrical distribution to the site infrastructure will consist of a dedicated 27.6 kV overhead line distribution network, equipped with 4/0 ACSR conductors.

Fresh water will be obtained from the West Creek Pond and will be used for reagent preparation, surface utilities and for dust suppression. The total fresh water requirement is estimated to be approximately 75 m³/h.

Geotechnical

AMEC carried out a series of geotechnical drilling campaigns at the open pit, tailings and water dam sites, mineral waste stockpiles and critical areas between the open pit and Pinewood River to characterize the site conditions and the subsurface stratigraphy, and to determine the soil and rock characteristics relevant to design of the facilities.

The geotechnical site investigations for the purposes of slope and dam design included 16 boreholes for the tailings dams, five (5) boreholes for the overburden stockpile and five (5) boreholes at the mine rock stockpile. An additional 26 boreholes were drilled to allow design of the open pit overburden slopes, obtain samples for characterization of the mine waste overburden for use as dam construction material, and to determine hydrolgeologic conditions between the pit and the Pinewood River.

The geotechnical investigations for the process plant, carried out under BBA’s specifications and requirements, included a comprehensive drilling and bedrock depth probing program consisting of two (2) drilling campaigns comprising 34 geotechnical boreholes, 38 test pits and 73 dynamic cone penetration tests. An iterative process between AMEC and BBA was followed in order to locate the process plant facilities on bedrock. The foundation design recommendations for the plant facilities were incorporated into the design of the facilities by BBA.

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1.18.1

Site Water Management

The water management system developed by AMEC is designed to generate a reliable water source for mill operations and ancillary uses, while optimizing the quantity and quality of site effluents released to the environment. Water will be recycled from various man-made ponds for mill process water in order to minimize the volume of fresh water to be taken from local watercourses. The system has been designed to ensure a reliable water supply at all times of the year and to allow for contingencies, such as dry years.

The system includes five (5) constructed ponds for water management in addition to sediment control ponds and a primary freshwater source. A constructed wetland is proposed downstream of the TMA and may act as part of the site effluent treatment system.

1.18.2

Tailings Management Area

The TMA has been designed by AMEC to store the process plant tailings. The tailings are assumed to be Potentially Acid Generating (“PAG”) and are therefore deposited in an area which allows maintenance of a permanent water cover for closure. The total volume of tailings produced over the mine life will be approximately 82 Mm³ at a deposited dry density of 1.4 t/m³.

The TMA location to the northwest of the open pit was selected in consideration of the topography, location of the pit and watershed boundaries, availability of dam construction materials and suitability for a flooded water cover for closure.

The geometry of the dams allows for a significant portion of the construction using haul trucks from the mine fleet. The relatively flat dam slopes required are constructed using overburden placed by the mine fleet, while the dam core, filter, drain and erosion protection zones constructed by a qualified earthworks contractor.

1.19

Market Studies and Contracts

Neither BBA, nor Rainy River, has conducted a market study in relation to the gold and silver doré which will be produced by the Rainy River Gold Project. Gold and silver are freely

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traded commodities on the world market for which there is a steady demand from numerous buyers.

There are no refining agreements or sales contracts currently in place that are relevant to this Technical Report.

1.20

Environmental and Permitting

Environmental aspects have figured prominently in the development of the preliminary layouts and designs for the Rainy River Gold Project described in this report. These include consideration of the implications of design alternatives from an environmental management and approvals perspective, related to mineral waste management, and the siting and location of facilities and infrastructure. From an environmental perspective, the Rainy River Gold Project is unique in that there are no lakes located within, or adjacent to, the main site. Additionally, the creeks and streams that are present do not support a commercial or recreational fishery.

There is considerable environmental baseline information currently available regarding the site and the surrounding area, compiled through extensive field investigations conducted over a four-year period. This information is being augmented as appropriate to support the progressing engineering design. Based on the information available to date and our understanding of the proposed development, there are no environmental aspects that are considered limiting to the Project development.

Most mining projects in Canada are reviewed under one (1) or more Environmental Assessment (“EA”) and approvals processes. The Rainy River Gold Project has initiated a coordinated Federal EA/Provincial Individual EA. A small number of Federal environmental approvals are likely necessary, as well as an anticipated requirement for a Schedule 2 listing for mineral waste

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management. A number of Provincial environmental approvals will also be needed, focussed on water taking, effluent and emission management, and closure planning.

The objective of final reclamation for the Rainy River Gold Project is to return the site to a productive condition on completion of mining activities. A closure plan must be filed and financial assurance provided to the Province before construction of the Project is initiated. A preliminary closure approach that is consistent with regulatory requirements has been developed for the Rainy River Gold Project. As possible, work will be completed progressively during operations; an industry best management practice. The financial model for the Project includes consideration of these costs.

1.21

Capital and Operating Costs

Capital Costs

The Capital Cost Estimate for the Project was developed by BBA, with input from various consultants according to their scope of work. The Capital Cost Estimate is based on a combination of equipment supplier quotes, supplier pricing, construction contractor input and experience with similar sized operations. This Project estimate meets AACE Class 3 requirements and is prepared to form the basis for budget authorization, appropriation and/or funding purposes. It has an expected accuracy range of -10 %/+15 %. This capital cost estimate assumes contracts will be awarded to reputable contractors on a lump sum basis and an open shop environment.

The projected pre-production capital cost for the Project is estimated to be $713M, including a $55M contingency allocation (based on 96% probability that the Project will be less than or equal to the stated Project cost). Total sustaining capital costs for the open pit mine are $322M, while underground mine development and sustaining capital costs are estimated to be $68M and $95M, respectively. Working capital requirements, to cover the period between commission and first metal sales, and project closure bonding are assumed to be covered by the Project financing or a negotiated loan facility and are not included in the cost estimate. The Project capital summary is outlined in Table 1-12, as follows:

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Table 1-12: Overall Project Capital Cost Summary


Area Description	Pre-Production Capital Costs ($M)		Sustaining Capital Costs ($M)
Overhead Power Line		11.1
Highway 600 Realignment		10.1
Open Pit Pre-Stripping		43.2		70.7
Open Pit Mine Equipment		26.2		161.5
Open Pit Waste Removal¹		45.1		62.7
Underground Mine – Development Capital²				67.8
Underground Mine Sustaining Capital³				94.6
Site Development		81.2		6.0
Process Facilities		283.7		1.5
Tailings and Water Management		44.9		23.6
Housing and Equipment Salvage Value				(64.2	)
Reclamation and Closure Costs				60.1
Indirect Costs		112.8

Subtotal		658.3		484.3

Contingency		55.0

Total Capital		713.3		484.3

¹	Capitalization of waste rock removal costs during production years with stripping ratios higher than 3.10 (LOM).

²	Funded through internal cash flows, this is the capital required in the development phase of the underground mine, consisting of equipment and infrastructure, as well as vertical and horizontal development.

³	Funded through internal cash flows, this is the sustaining capital required for the underground mine, consisting of equipment and infrastructure, as well as vertical and horizontal development.

Mining equipment quantities and costs have been developed based on the mine plan. Mining equipment costs are based on quotes requested during this study and from BBA’s recently updated database of supplier pricing. In order to reduce initial mining equipment costs, it is assumed that Rainy River will finance the equipment purchases with the equipment suppliers. The initial mining equipment CAPEX covers the initial deposit and the financing costs of all equipment required in pre-production.

Pre-stripping costs of overburden to remove an estimated 80M tonnes have been capitalized over a period of seven (7) years, including the first two (2) years of pre-production. The average unit rate used for pre-production stripping and overburden removal is $1.42/t, based on using the open pit mining equipment fleet.

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The site plan and General Arrangement (“GA”) drawings developed in this Study have been used to estimate quantities and generate Material Take-Offs (“MTOs”) for all commodities. Equipment costs have been estimated using budgetary proposals obtained from suppliers for all major process equipment. For process and mechanical equipment packages, equipment datasheets and summary specifications were prepared and budget pricing obtained from suppliers. For packages of low monetary value, pricing was obtained from BBA’s recent project database. A detailed equipment list was developed with equipment sizes, capacities, motor power, etc. Related infrastructure costs were estimated by BBA based on the site plan developed.

Underground mine development costs of $107.6M and equipment costs of $54.8M are included in the sustaining capital and were provided by Golder Associates.

During life of the operation, an estimated sustaining capital of $23.6M is required for necessary additions or improvements to the tailings management area. The main components of sustaining capital related to the TMA and water management include:

•

Phased construction of TMA dams based on the tailings management strategy developed by AMEC;

•

The construction of an artificial wetland; and

•

A tailings pump booster station and tailings pipeline expansion during operation.

Progressive rehabilitation and mine closure quantities have been estimated by AMEC. BBA has estimated unit costs for the material handling requirements based on local contractor estimates and the use of the RRR mine equipment.

1.22

Operating Costs

Introduction

The cost of electrical power was taken to be $0.060/kWh, which includes a $0.02/kWh reduction for the current Northern Ontario Industrial Rebate Program and other potential future government programs. The cost of propane and diesel were assumed to be $0.50/L and $0.85/L, respectively. Peak operations personnel will consist of approximately 601 persons, including 89 employees in the process plant, 26 employees in G&A, 299 employees in the open pit mine, 187 employees in

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the underground mine including eight (8) contracted employees. Project operating costs by area are summarized in Table 1-13.

Table 1-13: Key Project Operating Costs


Area	Units	Unit Costs (LOM)
Open Pit Mining (Waste + Ore + Stockpile)^1,2	$/t milled		7.19
Underground Mining³	$/t milled		2.02
Processing	$/t milled		8.65
General and Administrative	$/t milled		1.21
Refining and Transportation	$/t milled		0.14
Royalty Payments	$/t milled		0.54

Total	$/t milled		19.75

¹	Equivalent to $1.95/tonne mined.

²	Stockpile reclamation costs of $0.337 per tonne milled are included in the open pit mining costs.

³	Equivalent to $75.52/tonne mined.

Open Pit Mining

Open pit mining operating costs were developed for the first ten (10) years of production and include operating costs for the fleet of primary, support and auxiliary equipment, maintenance, electricity and fuel costs, hourly labour and salaried personnel, blasting costs and services. The mine mobile fleet unit operating cost is based on both suppliers data requested during this study and from BBA’s recently updated database. The LOM mining cost is $1.95/t mined.

Stockpile Reclaiming

Stockpile reclaiming operating costs were developed for the last six (6) years of production. The costs use the same basis as the mine operating costs and represent the primary and support fleet and labour required during the reclamation of the low grade ore stockpile. During the stockpile reclamation period, a total of 7.7 Mt of ore per year at an average rehandling cost of $0.91/tonne will be trucked to the primary crusher.

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Underground Mining

The operating costs for the Rainy River Underground mine (“RRU”) were estimated from first principles using a detailed cost model developed by Golder Associates. Supplier quotations were obtained for equipment operating costs and major consumables such as explosives, ground support, and cement. Labour costs are consistent with those used by BBA for the open pit mining and some adjustments were made for underground mining. Also included in the operating costs are:

•

Services such as setting up a drawpoint for remote mucking, definition diamond drilling and installing air and water pipes;

•

Cemented backfill of the primary and longitudinal stopes, and rockfill for the secondary and cut and fill stopes;

•

Operating development such as the excavation of the cross-cuts and drawpoints for the longhole stopes and attack ramps for the cut and fill stopes; and

•

Fixed costs such as electrical power for the main fans and pumps, and propane for the heaters.

Each item in the table contains the cost of consumables and equipment required to complete the activity. For example, the drilling activity contains the cost of rods, bits and adapters required to drill the hole and the cost to maintain the longhole drill and the cost of the fuel to move the longhole drill from one stope to the next.

The ore haulage costs include the trucking costs to haul ore out of the mine but the cost of the return trip is allocated either to ore haulage or backfill, depending on the destination of the returning truck. It is estimated that the majority of the development waste created will be used to backfill the cut and fill areas. The combined operating cost as shown in Table 1-14 for both underground mining methods is estimated to be $75.52/tonne mined.

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Table 1-14: Underground Mining Operating Costs¹


Item	Longhole Operating Cost ($/tonne mined)		Cut and Fill Operating Cost ($/tonne mined)
Drilling		1.52		1.32
Blasting		1.37		5.26
Mucking		2.78		1.31
Ground support		0.25		9.30
Services provision		0.69		3.32
Ore haulage (avg.)		1.56		2.02
Backfill		10.33		5.11
Operating development		5.07		10.50

Subtotal		23.57		38.14

Labour		37.73
Fixed (electrical power, propane)		11.91

Total		75.52

¹	The approximate tonnage breakdown is 76% longhole, 21% cut and fill and 3% development. All figures are rounded to reflect the relative accuracy of the estimate.

Process Plant

Process plant operating costs were calculated for 16 years of operation. The operating costs are based on metallurgical testwork, the mine plan, a recent salary survey, literature, and recent supplier quotations. The average life-of-mine processing operating costs were determined to be $8.65 per tonne milled at approximately 21,000 tpd. The yearly tonnages from the mine plan vary and were used to obtain the operating costs. As expected, operating costs for the first year are higher as the plant is still in ramp up. The operating cost includes reagents, consumables, grinding media, personnel (including contractors), electrical power, propane, and maintenance parts. The consumables accounted for in the operating cost include spare parts, grinding media and liner and screen components.

General and Administrative

General and Administrative (“G&A”) costs are expenses not directly related to the production of goods and encompass items not included in mining, processing, refining and transportation costs. These costs are based on the Owner’s recommendations, similar sized operations, and BBA’s in-house database.

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The G&A costs were calculated for 16 years of operation and are estimated to average $1.21 per tonne milled. This cost includes:

•

Human Resources;

•

Site Administration, Management and Insurance;

•

Infrastructure Power;

•

Health and Safety Supplies;

•

Assay Laboratory Supplies;

•

Environmental Costs;

•

G&A Personnel;

•

Information Technology (“IT”); and

•

Training.

Royalties, Refining and Transport

Annual royalty costs were provided by Rainy River and are based on the conceptual mine design and production profile, along with the terms of the individual royalty agreements. Refining and transportation costs are based on a quotation from a North American gold refinery.

1.22.1

Cash Costs

Table 1-15 provides a summary of average cash costs per ounce of gold over the first 10 years of the Project, as well as over the life-of-mine. Cash costs are calculated by recording the mining cost of stockpiled material in the periods in which the material is processed and revenue recognized, in accordance with IFRS. The gold cash costs, including silver credits and royalties total USD $468/oz. Au for the first ten (10) years of operation and USD $544/oz. Au over the life-of-mine. As noted in Section 1.17, waste rock costs during operating years with high stripping ratios (>3.10) have been capitalized based on the quantity of material above the average LOM stripping ratio.

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Table 1-15: Average Gold Cash Costs¹


Area	Initial 5 Years USD$/oz. Au			Initial 10 Years USD$/oz. Au			LOM USD$/oz. Au
Mining (Open Pit and Underground)²		225			268			275
Processing		184			188			258
General and Administrative		27			28			36
Refining Expenses		4			4			4
Royalties		6			18			16
Silver Credit		(33	)		(38	)		(45	)

Total		413			468			544

¹	Includes silver credit and royalty payments. Cash costs are calculated by recording the mining cost of stockpiled material in the periods in which the material is processed and revenue recognized, in accordance with the IFRS.

²	During years with high stripping ratios (>3.10), the operating costs associated with rock waste material have been capitalized.

The cash costs per ounce of gold vary significantly, depending on the feed grade, mine strip ratio and the amount of stockpiling. The annual fluctuation of operating costs per ounce of gold produced can be seen in Figure 1-6. It should be noted that the overall costs per ounce increase quite substantially during the later years due to the processing of stockpile material (with gold grades ranging from 0.3 to 0.6 g/t Au).

LOGO

Figure 1-6: Annual Operating Cash Costs (USD/oz. Au) with Silver Credit

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1.23

Economic Analysis

A financial analysis for the Rainy River Gold Project was carried out using a discounted cash flow approach. The internal rate of return (“IRR”) on total investment was calculated based on 100% equity financing even though RRR may decide in the future to finance part of the Project with debt financing, equipment financing and/or a streaming agreement. The Net Present Value (“NPV”) was calculated from the cash flow generated by the project based on a discount rate of 5%. The payback period based on the undiscounted annual cash flow of the project was also indicated as a financial measure. The Financial Analysis was performed using the following assumptions and basis:

•

The base case gold and silver prices are USD $1,400/oz. and USD $25/oz., respectively;

•

Commercial production will begin in Q3 of 2016;

•

The United States to Canadian dollar exchange rate has been assumed to be USD $1.00:CAD $1.00 during the first two (2) preproduction years and USD $1.00:CAD $1.07 during operations;

•

All cost and sales estimates are in constant Q4 2012 Canadian dollars with no inflation or escalation factors taken into account;

•

All gold and silver is sold in the same year of production;

•

All project related payment and disbursements incurred prior to the effective date of this report are considered as sunk costs. Disbursements projected for after the effective date of this report but before the start of construction are considered to take place in the pre-production period;

•

The after tax economics were provided by RRR and an external tax consultant. BBA has not verified this work; and

•

All values shown are post payment of royalties.

The results of the economic analysis represent forward-looking information and are subject to a number of known and unknown risks, uncertainties, and other factors that might cause actual results to differ materially from those presented here. The base case general economic results are summarized in Table 1-16.

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Table 1-16: Financial Analysis Summary


Description	Base Case		Units
Average Daily Milling Rate		tpd		21,000
Open Pit Mining (waste, ore and stockpiling)¹		$/t milled (LOM)		7.19
Stockpile Reclaim		$/t milled (LOM)		0.35
Underground Mining²		$/t milled (LOM)		2.02
Processing		$/t milled (LOM)		8.65
General and Administration		$/t milled (LOM)		1.21
Refining and Transportation Expenses		$/t milled (LOM)		0.14
Royalties		$/t milled (LOM)		0.54
Gold Recovery		% (LOM)		90.4
Silver Recovery		% (LOM)		64.1
Initial Project Capital Cost		$M		713.3
Open Pit Sustaining Capital Cost (from internal cash flows)		$M		321.9
Underground Development Capital Cost (from internal cash flows)		$M		67.8
Underground Sustaining Capital Cost (from internal cash flows)		$M		94.6
Five (5) year Au Cash Cost³		USD/oz.		413
Ten (10) year Au Cash Cost³		USD/oz.		468
LOM Au Cash Cost³		USD/oz.		544
Net Present Value (5% disc) (After-Tax)		$M		931
Internal Rate of Return (After Tax)		%		23.7
Simple Payback Period (After Tax)		Years		3.2

¹	Equivalent to $1.95/tonne mined.

²	Equivalent to $75.52/tonne mined.

³	Cash costs include silver credits and royalty payments. Cash costs are calculated by recording the mining cost of stockpiled material in the periods in which the material is processed and revenue recognized, in accordance with the IFRS.

As part of the financial results, the Life-of-Mine Cash Flow Projection has been calculated and is shown in Figure 1-7. The graph includes cumulative cash flow projections (non-discounted) as well as discounted (5%) cumulative cash flow projections on a pre and post-tax basis.

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LOGO

Figure 1-7: Life-of-Mine Cash Flow Projections

An after-tax sensitivity analysis was also performed to ascertain the impact of changes in metal price, capital costs, operating costs and foreign exchange rates. The results, for both the Net Present Value using a discount rate of 5% and the Internal Rate of Return sensitivity analysis, are shown in Figure 1-8 and Figure 1-9, respectively.

LOGO

Figure 1-8: After-Tax Net Present Value (NPV) Sensitivity Analysis at 5% Discount Rate

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LOGO

Figure 1-9: After-Tax Internal Rate of Return (IRR) Sensitivity Analysis

1.24

Adjacent Properties

Seven (7) properties in the exploration stage are located adjacent to or near the Rainy River Gold Project property. Bayfield Ventures Corp. holds three (3) of the properties, known as the B Block, C Block and Burns Block. In 2012, Rainy River also purchased a 100% interest in the surface rights to the Burns Block. Although several significant gold mineralized intersections on the Burns Block have been reported by Bayfield, to Rainy River’s knowledge the Burns Block does not contain any “mineral resource” (as that term is defined by the Canadian Institute of Mining, Metallurgy and Petroleum, and incorporated by reference in National Instrument 43-101), nor has there been a 43-101 compliant technical report completed for the Burns Block. Coventry Resources Inc. holds three (3) of the properties, known as the Pattullo, Nelles and Blue properties. King’s Bay Gold Corp. holds the seventh property, being a single continuous land package contiguous with the most northerly portion of the Rainy River Gold Project property. The closest Canadian operating mine is the Lac des Iles; palladium, nickel, gold and copper mine, owned by North American Palladium Ltd., located 90 km northwest of Thunder Bay and nearly 400 km northeast of the Rainy River Gold Project.

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1.25

Other Relevant Data and Information

Project Development Schedule

A Project Development Schedule has been generated to achieve production in Q3 of 2016. The schedule includes consideration of early work requirements, various studies, the Environmental Assessment process, engineering, procurement, and construction activities. It covers all required site infrastructures, crushing, mineral resource stockpiling, milling and tailings management. The preliminary on-site personnel requirement peaks at approximately 405 persons during the construction of the Project. The major Project activity durations and milestones are listed below in Table 1-17.

Table 1-17: Rainy River Project Development Activities


Activity	Start Date	Completion Date
Testwork and Engineering Studies	Ongoing	Q2 2013
Detailed Engineering	Q3 2013	Q1 2015
Environmental Assessment, Receipt of Construction Permits	Q1 2013	Q3 2014
Highway 600 Road and Realignment	Q3 2014	Q1 2015¹
Power Line Construction and Activation	Q3 2014	Q1 2015
Pre-Stripping and Bench Preparation	Q3 2014	Q2 2016
Construction of plant site & infrastructure	Q3 2014	Q2 2016
Commissioning and Production	Q3 2016	Q4 2016

¹	A target date has been set for Q4 2014, while a more conservative guidance is provided above.

Some of the most time-sensitive items within the scope of executing the Rainy River Project are: environmental permitting, realignment of Highway 600, electrical transmission line permitting, integration to the electrical grid and construction of the process plant.

The Project execution will be managed by an EPCM (engineering, procurement and construction management) contractor that will be contracted out to qualified firms, under the supervision of the Rainy River Resources Project team.

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1.26

Conclusions

The Feasibility Study indicates that the Rainy River Gold Project, based on the calculated Proven and Probable reserves (shown in Table 1-18) of 116.3 Mt grading 1.08 g/t Au and 2.76 g/t Ag, can support a 20,000 tpd open pit and 1,000 tpd underground mine.

Table 1-18: Proven and Probable Reserves (April 10, 2013)


Resources Category	Tonnage (Mt)		Au Grade (g/t)		Ag Grade (g/t)		Au (In-Situ oz.)		Ag (In-Situ oz.)
Proven		27.7		1.14		1.94		1,014,584		1,727,979
Probable		88.6		1.06		3.01		3,016,924		8,587,034

TOTAL (Combined)		116.3		1.08		2.76		4,031,508		10,315,013

Mineralized material will be sent to a process plant designed to achieve gold and silver recoveries of 90.4% and 64.1%, respectively. It is anticipated that, over a mine life of 16 years, approximately 3,645 koz. of gold and 6,615 koz. of silver will be produced.

The initial capital cost of the Project is estimated to be $713M and the sustaining capital (including Underground mining) is estimated to be $484M. The life-of-mine total cost is USD $544 /oz. Au, including silver credits and royalty payments. The Project NPV (after-tax) is estimated to be $931M using a discount rate of 5%. The Project internal rate of return (after-tax) is estimated at 23.7% and the simple payback period (after-tax) is 3.2 years.

The after-tax sensitivity analyses indicates that positive Project returns can be achieved over the likely range of variation in gold prices (± 20%), metal recovery (- 10% / +3%), capital costs (± 20%) and operating costs (± 20%).

In BBA’s opinion, the Rainy River Gold Project is sufficiently robust to warrant advancing to the next stage of development, that being the start of detailed engineering and construction.

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1.27

Recommendations and Future Work Program

The following recommendations are made considering the results of the Feasibility Study and the Project risks identified. BBA recommends a work program that includes further condemnation drilling, metallurgical testing and various studies, aiming at completing the characterization of the Project in preparation for the detailed engineering study phase. The suggested work program includes the following components:

•

Continuation of open pit and underground mine design optimization;

•

Procurement of long lead time mining and process equipment;

•

Continuation of preparatory work to secure electrical power and procurement of long lead time electrical equipment;

•

Continue recruiting key personnel;

•

Continuation of environmental assessment studies;

•

Continuation of First Nation and public consultations;

•

Continuation of hydrogeological studies in specific areas;

•

Secure Project financing;

•

Secure required permits and authorizations from government and regulatory agencies;

•

Initiate detailed engineering activities; and

•

Construction of the Project (following appropriate approvals).

The costs of this next engineering phase and additional exploration activities are estimated at approximately CAD $39.6M, as shown in Table 1-19.

Table 1-19: Budget for 2013


Activity	Cost ($ M)
Exploration		18.9
Condemnation Drilling		2.7
Basic and Detailed Engineering Activities		18.0

Total		39.6

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2.	INTRODUCTION

The Rainy River Gold Project is an advanced stage gold exploration project situated in the southern half of Richardson Township, approximately 50 km northwest of Fort Frances in northwestern Ontario, Canada. Richardson Township is one of several townships in the area that comprises Chapple Township, the municipal organization in the area. The Project is located in Richardson Township, approximately 162 km south of Kenora and 418 km west of Thunder Bay. In June 2005, Rainy River acquired the Project from Nuinsco and currently holds a 100% interest in the Project.

2.1

Scope of Study

The following Technical Report (the “Report”) presents the results of the Feasibility Study for the development of the Rainy River Gold Project, in Northwestern Ontario. In Q3 of 2011, Rainy River commissioned the engineering consulting group, BBA, to lead and perform the Study, based on contributions from a number of independent consulting firms. This Report was prepared at the request of Mr. Garett Macdonald, Vice President, Operations of Rainy River Resources. As of the date of this Report, Rainy River Resources is a Canadian publicly traded company listed on the Toronto Stock Exchange (“TSX”) under the trading symbol RR, with its head office situated at:

1 Richmond Street West, Suite 701

Toronto, Ontario

Canada M5H 3W4

Tel: (416) 645-7280

This Report, titled “Feasibility Study of the Rainy River Gold Project, Ontario, Canada”, was prepared by Qualified Persons following the guidelines of the NI 43-101, and in conformity with the guidelines of the Canadian Institute of Mining, Metallurgy and Petroleum (“CIM”) Standards on Mineral Resources and Reserves.

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A summary of the Report contributors and their area of responsibility are presented in Table 2-1.

Table 2-1: Major Study Contributors


Consulting Firm or Entity		Area of Responsibility

SRK Consulting (Canada) Inc.		Geological modelling and resource definition.

BBA Inc.		Open pit mine design, processing plant, site infrastructure, capital costs, operating costs, financial analysis and overall integration.

Golder Associates Ltd.		Underground mine design, capital costs and operating costs.

AMEC Environment & Infrastructure		Tailings, waste rock and water management, closure plan and geotechnical studies.

2.2

Effective Dates and Declaration

This Report is in support of the Rainy River Resources Ltd. press release, dated April 10, 2013, entitled “Rainy River Resources Completes Feasibility Study: Establishes Intermediate Production Profile with 4.0 Million Ounces of Gold in Proven and Probable Reserves”. This report is considered effective as at April 10, 2013. BBA’s opinion contained herein is based on information collected by BBA throughout the course of BBA’s investigations, which in turn reflects various technical and economic conditions at the time of writing. Given the nature of the mining business, these conditions can change significantly over relatively short periods of time. Consequently, actual results may be significantly more or less favourable.

Pursuant to a takeover bid commenced on June 18, 2013, New Gold Inc. (“New Gold”) has acqured majority ownership of RRR. RRR continues to exist as a separate legal entity from New Gold. New Gold has requested that BBA readdress this Technical Report to New Gold in order to support its own disclosure. No changes have been made to this Technical Report beyond addressing it to New Gold, inserting references to New Gold where appropriate and re-dating it, as well as formatting changes and minor typographical corrections.

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This Report may include technical information which requires subsequent calculations to derive subtotals, totals and weighted averages. Such calculations inherently involve a degree of rounding and, consequently, introduce a margin of error. Where this occurs, BBA does not consider it to be material. The overall Study was collated and integrated by BBA personnel.

BBA is not an insider, associate or an affiliate of Rainy River and neither BBA nor any affiliate has acted as Advisor to Rainy River, its subsidiaries or its affiliates, in connection with this Project. The results of the technical review by BBA are not dependent on any prior agreements concerning the conclusions to be reached, nor are there any undisclosed understandings concerning any future business dealings.

This Report is intended to be used by Rainy River, subject to the terms and conditions of its agreement with BBA. This agreement permits Rainy River Resources to file this Report as an NI 43-101 Technical Report pursuant to provincial securities legislation. With the exception of the purposes legislated under provincial securities laws, any other use of this Report, by any third party, is at that party’s sole risk.

2.3

Sources of Information

This Report is based in part on internal company reports, maps, published government reports, company letters and memoranda, and public information, as listed in Section 27 “References” of this Report. Sections from reports authored by other consultants may have been directly quoted or summarized in this Report, and are so indicated, where appropriate.

It should be noted that the authors have made use of selected portions or updated excerpts from material contained in the following NI 43-101 Compliant Technical Report. “Preliminary Economic Assessment Update for the Rainy River Gold Project”, for Rainy River Resources Ltd. NI 43-101 Technical Report, prepared by BBA Inc., dated October 12, 2012. This Report is publicly available on SEDAR (www.sedar.com).

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This Feasibility Study has been completed using the previously mentioned Technical Report as well as available information contained in, but not limited to, the following reports, documents and discussions:

•

Technical discussions with Rainy River Resources personnel;

•

Personal inspection of the Rainy River Gold Project property;

•

Report of mineralogical, metallurgical and grindability characteristics of the Rainy River Gold deposit, conducted by SGS Minerals Services and other firms on behalf of Rainy River Resources;

•

Resource Block Model provided by SRK;

•

A conceptual process flowsheet developed by BBA based on the testwork and similar operations;

•

Internal and commercially available databases and cost models;

•

Various reports produced by AMEC concerning environmental considerations for permitting, site hydrology, hydrogeology and geotechnical, tailings management and site closure plan;

•

A memo from Merit Consultants Inc., providing construction labour rates and productivity factors;

•

A memo from SanZoe Consulting Inc., providing future electrical costs for industrial users in Ontario;

•

A Feasibility Study Report provided by TBT Engineering Ltd., to Rainy River Resources for the Highway 600 Realignment;

•

“Amended Technical Report for the Rainy River Gold Project, Northwestern Ontario, Canada”, for Rainy River Resources Ltd. NI 43-101 Technical Report, prepared by SRK Consulting (Canada) Inc., dated June 4, 2012 (Dorota El-Rassi, P.Eng. and Glen Cole, P.Geo.);

•

Internal unpublished reports received from Rainy River staff; and

•

Additional information from public domain sources.

BBA believes that the basic assumptions contained in the information above are factual and accurate, and that the interpretations are reasonable. BBA has relied on this data and has no reason to believe that any material facts have been withheld. BBA also has no reason to doubt the reliability of the information used to evaluate the mineral resources presented herein.

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2.4

Terms of Reference

Unless otherwise stated:

•

All units of measurement in the Report are in the metric system;

•

All currency amounts in this Report are stated in Canadian dollars (“CAD”), unless otherwise stated;

•

All ounce units are reported in troy ounces, unless otherwise stated; 1 oz. (troy) = 31.1 g = 1.1 oz. (imperial);

•

All metal prices are expressed in terms of US dollars (“USD”);

•

A foreign exchange rate of USD $1.00 = CAD $1.00 was used for the pre-production period and USD $1.00 = CAD $1.07 for the balance of the Project life; and

•

All cost estimates have a base date of the third (“Q3”) quarter of 2012.

Grid coordinates for the Block Model are given in the UTM NAD 27 (Zone 19N) and latitude/longitude system; maps are either in UTM coordinates or in the latitude/longitude system.

2.5

Site Visit

BBA and Rainy River Resource representatives conducted a site visit on October 11, 2012 and BBA was represented by Mr. David Runnels and Mr. Patrice Live. The purpose of the visit was to provide the Project team members with an overview of the Rainy River Gold Project property and to review Project development milestones and planning. Rainy River Resources geologists were available to discuss general geological conditions and to provide a tour of the core storage facility, with a presentation of select mineral material. To provide an overview of the Rainy River Gold Project property terrain and potential locations for eventual Project infrastructure, Mr. Garett Macdonald, of Rainy River Resources, led a visit of the Rainy River Gold Project property using a company vehicle. All qualified persons, with the exception of Mr. Colin Hardie, Mr. Glen Cole and Mr. Adam Coulson were present. Mr. Hardie visited the site in June 2011; Mr. Cole was present at the site from April 30 to May 2, 2013, and Mr. Coulson visited the site in January 2012.

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2.6

Acknowledgement

BBA would like to acknowledge the support provided by Rainy River Resources personnel during this assignment. Their collaboration is greatly appreciated. The Project also benefitted from the inputs of Mr. Kerry Sparkes, Vice President, Exploration, Mr. Garett Macdonald, Vice President, Operations, Mr. Michael Mutchler, Vice President & Chief Operating Officer, Mr. Paolo Toscano, Director of Metallurgy and Mr. Nick Nikolakakis, Vice President & Chief Financial Officer. Their contributions are gratefully acknowledged.

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3.	RELIANCE ON OTHER EXPERTS

BBA prepared this Feasibility Study using the reports and documents noted in Section 27 “References”. BBA has not performed an independent verification of land title and tenure as summarized in Section 4.1 of this Technical Report. BBA did not verify the legality of any underlying agreement(s) that may exist concerning the permits or other agreement(s) between third parties, but has relied upon the opinion of the client’s layers, Heenan Blaikie of Toronto, for the land tenure information summarized in Section 4. An extract from the land title opinion provided by Heenan Blaikie may be found in Appendix A. Any statements and opinions expressed in this document are given in good faith and in the belief that such statements and opinions are not false and misleading at the date of this Report. BBA is not aware of any known litigations potentially affecting the Rainy River Gold Project.

It should be understood that the mineral reserves and resources presented in this Report are estimates of the size and grade of the deposits. The estimates are based on a certain number of drill holes and samples, and on assumptions and parameters currently available. The level of confidence in the estimates depends upon a number of uncertainties. These uncertainties include, but are not limited to: future changes in metal prices and/or production costs, differences in size, grade and recovery rates from those expected, and changes in Project parameters. In addition, there is no assurance that the Project implementation will be carried out. BBA’s responsibility was to assure that this Technical Report met the stipulated guidelines and standards, given that certain sections of this Report were contributed by SRK, Golder, AMEC or other Rainy River Resources Ltd. consultants.

3.1

Report Responsibility and Qualified Persons

Table 3-1 outlines responsibility for the various sections of the Report and the name of the corresponding Qualified Person.

The authors of this Report consist of both BBA employees and various other consultants. Each author is a Qualified Person and is responsible for various sections of this Report, according to his or her expertise and scope of work. Each author has contributed figures, tables and portions of Sections 1 (Summary), 25 (Interpretation and Conclusions), and 26 (Recommendations).

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Table 3-1: Qualified Persons and Areas of Report Responsibility


Section	Description	Responsibility	Qualified Person	Comments and Exceptions
1.	Summary	BBA	C. Hardie	All contributed based on their expertise and scope of work.
2.	Introduction	BBA	C. Hardie
3.	Reliance on other Experts	BBA	C. Hardie
4.	Project Property Description and Location	BBA AMEC E&I RRR	C. Hardie	Mineral tenure and underlying agreements by RRR (Sections 4.1 and 4.2) and environment considerations and mining rights (Sections 4.3 and 4.4) by AMEC.
5.	Accessibility, Climate, Local Resource, Infrastructure and Physiography	SRK	G. Cole
6.	History	SRK	G. Cole
7.	Geological Setting and Mineralization	SRK	G. Cole
8.	Deposit Types	SRK	G. Cole
9.	Exploration	SRK	G. Cole
10.	Drilling	SRK	G. Cole
11.	Sample Preparation, Analyses and Security	SRK	G. Cole
12.	Data Verification	SRK	G. Cole
13.	Mineral Processing and Metallurgical Testing	BBA	D. Runnels	Testwork by SGS, FLS, Metso, SAGDesign and Outotec.
14.	Mineral Resource Estimate	SRK	D. El-Rassi
15.	Mineral Reserve Estimates	BBA	P. Live D. Tolfree	Underground mineral resources estimate by Golder (Section 15.2)

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Section	Description	Responsibility	Qualified Person	Comments and Exceptions
16.	Mining Methods	BBA Golder	P. Live D. Tolfree A. Coulson	Open pit mining and reserves by BBA, underground mining and reserves by Golder (Section 16.3). Geomechanical designs by AMEC (Section 16.2.1. and 16.3.1)
17.	Recovery Methods	BBA	D. Runnels
18.	Project Infrastructure	BBA AMEC E & I	D. Runnels D. Ritchie	Site infrastructure by BBA, tailings management area design by AMEC (Section 18.10) and Highway 600 realignment study conducted by TBT Engineering Consulting Group (Section 18.1.1). Geotechnical by AMEC (Section 18.2).
19.	Market Studies and Contracts	BBA	C. Hardie	No market study performed.
20.	Environmental Studies, Permitting, and Social or Community Impact	AMEC E & I	S. Daniel
21.	Capital and Operating Costs	BBA Golder	C. Hardie D. Tolfree	AMEC provided tailings area construction quantities and site closure plan, Merit Consultants provided construction labour rates and productivity factors, Golder provided CAPEX and OPEX for underground mining (Sections 21.5 and 21.15.4). RRR provided royalty costs and taxation information.
22.	Economic Analysis	BBA	C. Hardie	AMEC provided closure costs.
23.	Adjacent Properties	BBA	C. Hardie
24.	Other Relevant Data and Information	BBA	D. Runnels	Schedule developed by BBA, Merit Consultants, AMEC provided permitting information.

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Section	Description	Responsibility	Qualified Person	Comments and Exceptions
25.	Interpretation and Conclusions	BBA	C. Hardie	All contributed based on their expertise and scope of work.
26.	Recommendations	BBA	C. Hardie	All contributed based on their expertise and scope of work.
27.	References	BBA	C. Hardie

The following Qualified Persons (QPs) have contributed to the writing of this Report and have provided QP certificates, included at the beginning of this Report. All QPs have visited the Rainy River Gold Project property. The information contained in the certificates outline the sections in this Report that each of the QPs is responsible for.


• Colin Hardie, P.Eng.		BBA Inc.

• David Runnels, Eng.		BBA Inc.

• Patrice Live, Eng.		BBA Inc.

• Sheila Daniel, M.Sc., P.Geo.		AMEC Earth & Infrastructure

• David Ritchie, P.Eng.		AMEC Earth & Infrastructure

• Adam Coulson, PhD., P.Eng.		AMEC Earth & Infrastructure

• Dorota El-Rassi, P.Eng.		SRK Consulting (Canada) Inc.

• Glen Cole, P.Geo.		SRK Consulting (Canada) Inc.

• Donald Tolfree, P.Eng.		Golder Associates Ltd.

3.2

Other Study Contributors

The individuals listed below have contributed to the Feasibility Study and to this Report, and have extensive experience in the mining and metals industry or in a supporting capacity in the industry. They are not considered as QPs for the purpose of this NI 43-101 Report.


• David Sprott, P.Eng.	Underground Mining Engineering	Golder Associates Ltd.

• Jay Collins, P.Eng.	Construction Management	Merit Consultants Inc.

• Wayne Clark, P.Eng.	Electrical Transmission Line	SanZoe Consulting Inc.

• Rob Frenette, P.Eng.	Road Construction	TBT Engineering Consulting Group

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• Stuart McTavish	Metallurgy	SF McTavish Consulting

• Paolo Toscano	Metallurgy	Rainy River Resources Ltd.

• Mike Mutchler	Mining Engineering & Economics	Rainy River Resources Ltd.

• Garett MacDonald	Mining Engineering	Rainy River Resources Ltd.

Jay Collins, P.Eng., President of Merit Consultants Inc., provided construction labour rates and productivity factors. Wayne Clark, of SanZoe Consulting Inc., provided consulting for applications to IESO and Hydro One Network for the energy costs and advice for the high voltage power transmission line.

Drill core samples for metallurgical testing were collected and prepared by Rainy River Resources and submitted to SGS Minerals Services (Lakefield, Ontario, Canada), which is an accredited laboratory. Although BBA has reviewed the testwork results generated by SGS and believes that they are generally accurate, BBA is relying on SGS as an independent expert.

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4.	PROPERTY DESCRIPTION AND LOCATION

The Rainy River Gold Project is located approximately 50 km to the northwest of Fort Frances, the nearest large town in Western Ontario; refer to Figure 4-1. The village of Emo is located approximately 25 km to the south on Highway 11.

The Rainy River Gold Project property is comprised of a portfolio of unpatented mining claims located in the townships of Fleming, Menary, Potts, Richardson, Senn, Sifton and Tait, and leasehold interest mining rights land claims, patented mining rights and surface rights land claims located in Mather, Potts, Richardson, Senn, Sifton and Tait townships. All of the unpatented mining claims, leasehold mining rights lands, and patented mining rights and surface rights lands, within the above-mentioned townships, can be accessed by a network of secondary all-weather roads that branch off the well-maintained Trans-Canada Highways 11 and 71.

4.1

Mineral Tenure

The Rainy River Gold Project is comprised of a total of 232 patented mining rights and surface rights lands, unpatented mining claims and interest in three (3) leasehold interest mining rights claims. The Rainy River Gold Project property is collectively located in Fleming, Mather, Menary, Potts, Richardson, Senn, Sifton, and Tait townships. The Project property covers an aggregate area of 16,697.06 hectares.

In Ontario, Crown lands are available to licensed prospectors for the purposes of mineral exploration. Claim staking is governed by the Ontario Mining Act and is administered through the Provincial Mining Recorder and Mining Lands offices of the Ministry of Northern Development and Mines (“MNDM”).

The unpatented mining claims are comprised of a multiple of 16 ha (40 acres) square blocks. In Ontario, after staking, the unpatented mining claims are recorded within 31 days with the MNDM upon payment of an appropriate fee. In order to keep the unpatented mining claims valid, an approved expenditure per claim in excess of CAD $400 within two (2) years is required.

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Patented lands do not have an assessment work obligation, but require maintaining both municipal realty and provincial mining land taxes.

In June 2005, Rainy River completed the acquisition of a 100% interest in the Project from Nuinsco Resources Limited (“Nuinsco”). As of March 11, 2013, the Rainy River Gold Project land package consists of:

•

A total of 151 patented mining rights and surface rights land claims, including three (3) leasehold interest patented mining rights land claims; and

•

A total of 81 unpatented mining claims.

An updated listing of the current Rainy River patented, leasehold and unpatented mining claims and leases was provided to BBA by Rainy River for review. These claims have not been legally surveyed by Rainy River.

Certain unpatented mining claims on a spot check basis were independently verified by BBA on May 15, 2013 by accessing the MNDM website:(www.mndmf.gov.on.ca/mines/claimaps_e.asp).

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LOGO

Figure 4-1: Location of Rainy River Gold Project (as of March 13, 2013)

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A plan illustrating the distribution of the claims on the Rainy River Gold Project is shown in Figure 4-1 and a title list is provided in Appendix B (claims as of March 13, 2013).

All unpatented mining claims are recorded in the name of Rainy River, save and except those unpatented mining claims set out within the ‘English Option’ and the ‘Roisin Option’ and the ‘Timberridge Option’, are described in more detail in Appendix B, and as of the effective date of this technical report are in good standing and have sufficient work assessment credits available for several years. BBA is not aware of any outstanding aboriginal land rights or aboriginal claims to this area.

The mineral resources reported herein occur within the patented claims former 4950, 5614, 5939, 25891, 25892 and 25894 (Figure 4-2).

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LOGO

Figure 4-2: Land Tenure Map of the Rainy River Gold Project (as of March 13, 2013)

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4.2

Underlying Agreements

Rainy River, through direct ownership or option agreements entitling it to purchase properties, has a 100% interest in the lands forming the Rainy River Gold Project.

Various option payments are due on certain patented land claims. Details of these option agreements, that include issuance of common shares and cash payments, are presented in Appendix B.

On March 3, 2010, Rainy River announced an option agreement to acquire five (5) unpatented mining claims in the Tait Township (Figure 4-2) from Rubicon Minerals Corporation in consideration of cash payments of CAD $150,000, and the issuance of 60,000 shares over a period of five (5) years and a 2% net smelter return royalty. Rainy River may, at any time, re-purchase one-half of the royalty for CAD $1,000,000.

On December 16, 2011, Rainy River entered into an option agreement to acquire three (3) unpatented mining claims in Richardson and Potts Townships (Figure 4-2) from Fred A. Roisin, in consideration of cash payments of CAD $100,000, and the issuance of 50,000 shares over a period of five (5) years and a 2% net smelter return royalty. Rainy River may, at any time, re-purchase one-half of the royalty for CAD $1,000,000.

On March 29, 2012, Rainy River entered into an option to acquire one (1) unpatented mining claim in Sifton Township (Figure 4-2) from Timberridge Land & Forestry Services Inc. in consideration of cash payments of CAD $100,000 and the issuance of 50,000 shares over a period of five (5) years and a 2% net smelter return royalty. Rainy River may, at any time, re-purchase one-half of the royalty for CAD $1,000,000.

4.3

Environmental Considerations

Rainy River initiated an environmental study of the Project area in 2008 by commissioning Klohn Crippen Berger Ltd. (“KCB”) to undertake a preliminary socio-environmental baseline and scoping study. Rainy River significantly expanded the scope of the baseline environmental assessment with KCB beginning in 2009 and again in 2010. In 2011, Rainy River commissioned AMEC

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Environment & Infrastructure (“AMEC”) to conduct further environmental baseline investigations from early 2011 to present, as well as anticipated mine environmental assessment and permitting stages. Five (5) years of comprehensive baseline field environmental study of the project and regional area have now been completed. From an environmental perspective, the Rainy River Gold Project is unique in that there are no lakes located within, or adjacent to, the Project footprint. Additionally, the creeks and streams that are present do not support commercial or recreational fisheries.

The objectives of the baseline studies are to characterize the natural (or biophysical) and human environment aspects of potentially impacted areas, along with reference locations (such as upstream locations) where appropriate for comparison. Environmental baseline data (description of the existing environment):

•

Helps inform project designs (for example, knowledge of rock characteristics assists in determining how best to handle and store the material);

•

Will allow an assessment to be made of likely project environmental effects, including comparisons with established environmental guidelines, thresholds and limits, where applicable; and

•

Provide a reference for future environmental monitoring (that is, it allows a comparison to be made of pre-development and post development conditions).

Baseline environmental studies undertaken include:

•

Air quality;

•

Meteorology and climate;

•

Sound;

•

Aquatic resources (fish and benthic invertebrates) and habitat;

•

Wildlife and habitat;

•

Species-at-Risk;

•

Surface water quality and flows;

•

Groundwater quality and paths;

•

Sediment quality;

•

Geochemistry;

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•

Socio-economics;

•

Archaeology and heritage resources; and

•

Traditional knowledge and traditional land use.

Rainy River and the seven (7) member First Nations of the Fort Frances Chiefs Secretariat (the “FFCS”) signed a Memorandum of Understanding (the “MOU”) in May 2010, which committed Rainy River to informing the FFCS in advance about exploration proposals and schedules, and conducting exploration activities in an environmentally responsible manner. Employment and contracting opportunities were also part of the terms including the joint initiative to fund the First Nations Engagement Specialist position for the FFCS to act as a liaison between Rainy River and the signatory First Nations. The MOU also committed Rainy River and the FFCS to developing and implementing a Participation Agreement (also known as an Impact and Benefits Agreement) that would include provisions for: mineral production support, consultation protocols, respect for Traditional Territories, training and employment, among other aspects. Pursuant to the MOU, Rainy River and the signatory First Nations held various community and leadership meetings which culminated in Rainy River and six member First Nations of the FFCS entering into a Participation Agreement in March of 2012 (the “PA”). Pursuant to the PA, a Participation Agreement Advisory Committee (“PAAC”) has been formed, comprised of one representative from each of the signatory First Nations and two (2) representatives from Rainy River. The PAAC currently meets on a regular basis to share information and pursue the successful implementation of the PA.

Rainy River signed an MOU on March 6, 2012 with the Big Grassy River First Nation. Rainy River is also engaged in discussions with Big Island First Nation, Onigaming First Nation, Naotkamegwanning (Whitefish Bay) First Nation, Buffalo Point First Nation and the Métis Nation of Ontario.

Rainy River has also kept the general public updated on Project plans. Community Town Hall meetings to discuss the Project are held regularly to help people understand the various stages of exploration as the Project moves towards a production decision. In addition, open houses have been held in various local communities to support the environmental assessment process.

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4.4

Mining Rights in Ontario

The Rainy River Gold Project is located in Ontario; a province that has a well understood permitting process in place and one that is coordinated with the Federal regulatory agencies. As is the case for similar mine developments in Canada, the Project will be subject to a Federal and Provincial Environmental Assessment (the “EA”) process. The Rainy River Gold Project will require completion of a Federal Environmental Assessment, pursuant to the Canadian Environmental Assessment Act, 2012. Rainy River received the Environmental Impact Statement Guidelines from the Canadian Environmental Assessment Agency on December 18, 2012. Rainy River entered into a Voluntary Agreement with the Ontario Ministry of the Environment to conduct a Provincial Individual Environmental Assessment that will meet the requirements of the Ontario Environmental Assessment Act. Several aspects of the Project were anticipated to require completion of Provincial Environmental Assessment process(es) and a single Provincial process coordinated with the required Federal Environmental Assessment process, was selected as the best approach to meet those (or other) needs. Terms of Reference for the Provincial Environmental Assessment were approved by the Ontario Ministry of the Environment on May 15, 2013. The same body of information will be used to inform the Provincial and Federal Environmental Assessment process, culminating in a single Environmental Assessment report that meets both the Federal Environmental Impact Statement Guidelines and the approved Provincial Terms of Reference.

After the EA processes are completed, environmental approvals will be required to construct, operate and close the Rainy River Gold Project. Due to the complexity and size of such projects, various Federal and Provincial agencies will have jurisdiction to either provide authorizations or permits that will enable project construction to proceed and to operate.

Federal agencies that will have significant regulatory involvement at the pre-production phase include the Canadian Environmental Assessment Agency, Environment Canada, Natural Resources Canada as well as Fisheries and Oceans Canada.

On the Provincial agency side, the Ministry of Northern Development and Mines (“MNDM”), Ministry of Natural Resources (“MNR”), Ministry of Environment (“MOE”), as well as the Ministry of Transportation (“MOT”) will each have key project development permit responsibilities.

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In support of the anticipated mine Environmental Assessment and associated permitting process, Rainy River has been in regular communication with each of the key regulatory agencies. Consultations with the Provincial and Federal regulatory agencies began in the spring of 2010. In support of the then anticipated Advanced Exploration permitting, Rainy River met with the pertinent agencies several times in 2010. With the recent decision to initiate Mine environmental assessment and permitting in 2012, Rainy River has held meetings with both the Canadian Environmental Assessment Agency as well as the Ontario Ministry of Environment.

The Project Environmental Assessment process will occur over an estimated 24-month period and will culminate with the issuance of various project authorizations (Federal) as well as operating permits (Provincial) over the ensuing approximate 12 months.

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5.	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

The Rainy River Gold Project is located approximately 50 km to the northwest of Fort Frances, the nearest large town in western Ontario. Infrastructure in the area of the Project is satisfactory with numerous gravel/paved roads, power and water resources available within close proximity to the Project. In addition, Rainy River Resources has adequate surface rights for the Project requirements, as proposed in the site plan (Appendix H).

5.1

Accessibility

The Project is centred in Richardson Township (part of Chapple Township) in northwestern Ontario (Figure 5-1), approximately 162 km (Highway 17/Highway 71/Regional Road 600) south of Kenora, and 418 km (Highway 11/Highway 71/Regional Road 600) west of Thunder Bay. These access roads are sealed allowing year-round access. The final access to the deposit consists of a 540 m track from Regional Road 600 to the 17 Zone and there are drill tracks to other areas, which are of exploration interest. Figure 4.2 in Section 4 indicates the location of the Project claims with respect to local roads in the area.

The Canadian National Railway is located 21 km to the south and runs east-west, immediately north of the Minnesota border. The nearby towns and villages of Fort Frances, Emo and Rainy River are located along this railway line.

5.2

Local Resources and Infrastructure

The towns within immediate driving distance of the Rainy River Gold Project are:

•

Emo, with a population of 1,305 – 34 km (30-minute drive);

•

Rainy River, population 909 – 57 km (80-minute drive); and

•

Fort Frances, with a population of 8,103 – 70 km (1-hour drive).

Hydroelectricity is produced north of Kenora at various locations, as well as west and east of Thunder Bay. A medium-sized coal-powered thermal power station is located east of Fort Frances and another is located near Thunder Bay.

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There is a ready supply of water in the area from lakes and rivers. Ground water is also likely to be in plenteous supply, given the abundance of standing water and rivers within the region. The major primary drainage system in the area includes Rainy Lake, which lies to the southeast and is drained by Rainy River which flows west along the Minnesota border to Lake of the Woods, which in turn feeds into the Lake Winnipeg watershed.

5.3

Climate

The climate is typically continental, with extremes in temperatures ranging from +35°C to -40°C, from summer to winter. Annual rainfall in the region averages about 60 cm, with heaviest rains expected from June to August, when an average of about 30 cm of rain is recorded. An average of 350 cm snowfall is recorded annually in the region.

5.4

Physiography

The Rainy River Gold Project region is divided into two (2) main physiographical regions. These regions are separated by a distinct northwest to southeast divider, locally termed the Rainy Lake - Lake of the Woods Moraine, which traverses the countryside immediately to the north of the Richardson Township. To the north and east of this Rainy Lake - Lake of the Woods Moraine, there is a substantial amount of bedrock exposure and topographic relief can be up to 90 m. This relief contrast is controlled by the geology of the batholiths, which erode negatively in comparison to the supracrustals of the Canadian Shield. The area was subjected to the Whiteshell glacial event from the Labradorean ice centre to the northeast.

The region to the south and west of the Rainy Lake - Lake of the Woods Moraine, comprises of lowlands, which underwent peneplanation in the Cretaceous, eroding away most of the Mesozoic cover. Topographic relief in this region is lacking, the glacial overburden is typically twenty to forty metres thick, drainage is poor and outcrop is limited to less than one percent of the surface area. This area was exposed to successive glaciations from the northeast and west.

The bedrock is immediately overlain by Labradorean till that is geochemically responsive. This Labradorean till is in turn overlain by thick, highly conductive glaciolacustrine silts and clays of

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Glacial Lake Agassiz and easterly transported clay and carbonate-rich Keewatin till. Some poorly drained areas are also covered by a thick peat layer which further impedes exploration activities.

The Rainy River Gold Project area is sparsely populated. The vegetation falls within the northeastern hardwood region immediately adjacent to the southern margin of the boreal forest (Figure 5-1).

LOGO

Figure 5-1: Typical Landscape in the Rainy River Gold Project Area

A:	Rainy River landscape looking south from the 433 Zone over the ODM and Beaver Pond Zones.

B:	View of the Rainy River Exploration Camp.

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6.	HISTORY

The bulk of this historical review is based upon the documentation of exploration in northwestern Ontario that is archived in the MNDM offices at Kenora. Exploration in the Rainy River Gold Project area began in 1967. Various companies were active between 1967 and 1989. Nuinsco undertook exploration activities between 1990 and 2004, with Rainy River continuing from 2005 onwards.

6.1

Previous Exploration Work

A summary of the exploration activities undertaken on the Rainy River Gold Project property from 1967 to 2004 is presented below.

6.1.1

Period 1967 to 1989 by Various Companies

1967

Anomalous copper was noted in the region.

1967

Noranda registered claims and performed geophysics.

1971

The Ontario Division of Mines, Ministry of Natural Resources, mapped the north-central part of the Rainy River Greenstone Belt (Blackburn, 1976).

1971

International Nickel Corporation of Canada (“INCO”) undertook follow-up ground geophysics. INCO drilled two (2) diamond drill holes in Richardson Township. Results are unknown.

1972

Hudson’s Bay Exploration and Development (“HBED”) undertook airborne and follow-up ground geophysics. In 1973, HBED drilled 54 core boreholes in the Rainy River Gold Project region. There was insufficient encouragement to continue and exploration was curtailed.

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1988

The Ontario Geological Survey (“OGS”) Map P.3140 was produced. It was based on the interpretation of aeromagnetic data and geological mapping carried out by Johns (1988). This mapping was supported by an OGS rotasonic drilling program on a 3 km drill grid completed between 1987 and 1988.

The OGS program resulted in the discovery of a “gold grains-in-till” anomaly in Richardson Township.

1988

Mingold Resources followed up on this gold grains-in-till anomaly and staked 85 claims and optioned patented lands in Richardson and some neighbouring townships. Their use of various sampling methodologies on the till, including reverse circulation drilling, gave inconclusive results.

6.1.2

Period 1990 to 2004 by Nuinsco

A tabulated summary of the exploration work undertaken by Nuinsco is provided in Appendix C.

1992

Nuinsco optioned patented lands centred on Richardson Township and the Menary Township. Gold occurrences discovered by King’s Bay Gold Corporation optioned from Western Troy Resources in September 1992.

1993 to 1998

A total of 597 widely spaced reverse circulation drill boreholes define a 15 km long gold grains-in-till dispersal train emanating from a 6 km2 “gold-in-bedrock” anomaly averaging 79 parts per million (“ppm”) gold.

1994 to 2004

Total of 217 core boreholes (49,515 m) drilled, mostly in Richardson Township.

1994

Investigation of gold grains-in-till and gold-in-bedrock anomalies with a series of core boreholes leading to the discovery of the 17 Zone.

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1995

Discovery of 34 Zone (copper-nickel-platinum group metals) followed by intensive diamond drilling.

1997

The discovery of 433 Zone 500 m to the north of the 17 Zone.

1999

Core drilling targeting the 34 Zone and a magnetic anomaly in Tait Township.

2000

Audio magneto-telluric (“AMT”) geophysical survey. Several targets were defined and drilled, but these proved to be massive graphite, disseminated sulphides or massive but barren sulphides.

2004

Investigation of the depth continuity of the 34 Zone with eight (8) core boreholes (1,549 m).

6.1.3

Previous Mineral Resource Estimates

Seven (7) previous mineral resource evaluations have been prepared for the Rainy River Gold Project by Mackie et al. in 2003, Caracle Creek International Consulting Inc. (“CCIC”) in 2008 and by SRK in 2009, 2010 and 2011. Five (5) of these mineral resource statements are documented in previous technical reports prepared for the Project and are available from SEDAR.

The initial mineral resource statement prepared by Mackie et al. (2003) is presented in Table 6-1. The second mineral resource statement was prepared by CCIC in 2008 and is presented in Table 6-2. CCIC reported the mineral resources at three (3) gold cut-off grades.

Table 6-1: Mineral Resource Statement* for the Rainy River Gold Project, Ontario,

Mackie et al., December 23, 2003


	Quantity		Grade								Metal
Category	‘000 t		Au g/t		Cu (%)		Zn (%)		Ag g/t		Au ‘000 oz.
Indicated		1,736		1.56		0.03		0.21		4.0		87.1
Inferred		11,025		1.33		0.02		0.20		3.6		471.2

*	Reported at a cut-off grade of 0.7 g/t gold.

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Table 6-2: Mineral Resource Statement for the Rainy River Gold Project, Ontario,

Caracle Creek International Consulting Inc., April 30, 2008


	Cut-off		Quantity		Grade				Metal
Category	Au g/t		‘000 t		Au g/t		Ag g/t		Au ‘000 oz.		Ag ‘000 oz.
Indicated		0.3		37,761		1.18		2.60		1,436		3,159
Inferred		0.3		79,654		0.94		2.31		2,400		5,923
Indicated		0.5		34,238		1.26		2.63		1,386		2,896
Inferred		0.5		67,564		1.03		2.35		2,233		5,109
Indicated		0.7		24,959		1.50		2.63		1,206		2,106
Inferred		0.7		44,391		1.25		2.26		1,787		3,257

In 2009, SRK prepared the third mineral resource statement incorporating information from an additional 112 core boreholes (59,719 m) drilled subsequent to the CCIC 2008 mineral resource statement. The third consolidated mineral resource statement prepared by SRK in April 2009 is presented in Table 6-3.

Table 6-3: Mineral Resource Statement* for the Rainy River Gold Project, Ontario,

SRK Consulting (Canada) Inc., April 28, 2009


	Quantity		Grade				Metal
Category	‘000 t		Au g/t		Ag g/t		Au ‘000 oz.		Ag ‘000 oz.
Open Pit**
Indicated
Volcanic-hosted
Within shell		55,027		1.21		1.89		2,135		3,350
Mafic-hosted
Within shell		57		1.37				3
Inferred
Volcanic-hosted
Within shell		12,491		0.95		2.36		382		950
Outside shell		50,637		0.79		2.19		1,278		3,562
Underground**

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	Quantity		Grade				Metal
Category	‘000 t		Au g/t		Ag g/t		Au ‘000 oz.		Ag ‘000 oz.
Indicated
Volcanic-hosted
Below shell		530		5.14		1.47		88		25
Inferred
Volcanic-hosted
Below shell		875		5.22		1.27		147		36
Combined Mining
Indicated		55,615		1.24		1.89		2,225		3,375
Inferred		64,003		0.88		2.21		1,807		4,548

Mineral resources are reported in relation to optimized pit shells. Mineral resources are not mineral reserves and do not have demonstrated economic viability. All figures are rounded to reflect the relative accuracy of the estimate. All assays have been capped where appropriate. SRK also estimated zinc, lead and copper grades but these metals are not reported in the mineral resource statement because they reasonably do not contribute to the metal value of the gold mineralization. Nickel, copper and platinum group metals were also estimated for one small zone (34 Zone, Domain 201) and are reported separately. The consolidated resource statement above includes the gold mineralization in 34 Zone.

Open pit mineral resources are reported at a cut-off grade of 0.4 g/t gold; underground mineral resources are reported at cut-off grade of 3.0 g/t gold. Cut-off grades are based on a gold price of USD $800 per ounce gold and a gold metallurgical recovery of 85%, without considering revenues from other metals.

In early 2010, SRK prepared the fourth mineral resource statement to incorporate information from 124 core boreholes (68,453 m) drilled on the Project during 2009. The fourth consolidated mineral resource statement dated February 26, 2010, is represented in Table 6-4.

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Table 6-4: Mineral Resource Statement*, Rainy River Gold Project, Ontario,

SRK Consulting (Canada) Inc., February 26, 2010


	Quantity		Grade				Metal
Category	‘000 t		Au g/t		Ag g/t		Au ‘000 oz.		Ag ‘000 oz.
Open Pit**
Indicated
Volcanic-hosted
Within shell		55,657		1.20		1.76		2,153		3,151
Mafic-hosted
Within shell		57		1.37		5.10		3		9
Inferred
Volcanic-hosted
Within shell		6,252		1.26		2.66		252		535
Outside shell		58,339		0.90		2.81		1,688		5,270
Underground**
Indicated
Volcanic-hosted
Below shell		1,119		6.03		4.28		217		154
Inferred
Volcanic-hosted
Below shell		4,339		5.15		1.69		718		236
Combined Mining
Indicated		56,833		1.30		1.81		2,370		3,314
Inferred		68,930		1.20		2.73		2,659		6,041

Open pit mineral resources are reported at a cut-off grade of 0.4 g/t gold, underground mineral resources are reported at a cut-off grade of 3.0 g/t gold. Cut-off grades are based on a gold price of USD $850 per ounce gold and a gold metallurgical recovery of 85%, without considering revenues from other metals.

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In early 2011, SRK prepared the fifth mineral resource statement to incorporate information from 163 core boreholes (84,648 m) drilled on the Project during 2010. The fifth consolidated mineral resource statement dated February 24, 2011 is represented in Table 6-5.

Table 6-5: Mineral Resource Statement*, Rainy River Gold Project, Ontario,

SRK Consulting (Canada) Inc., February 24, 2011


	Quantity		Grade				Metal
Category	‘000 t		Au g/t		Ag g/t		Au ‘000 oz.		Ag ‘000 oz.
Open Pit**
Measured		14,707		1.21		1.84		572		869
Indicated		77,934		1.05		2.24		2,640		5,616
Measured and Indicated		92,641		1.08		2.18		3,212		6,485
Inferred		104,591		0.80		2.31		2,703		7,781
Underground**
Measured		39		5.66		2.38		7		3
Indicated		1,197		5.18		3.30		199		127
Measured and Indicated		1,236		5.20		3.27		206		130
Inferred		3,831		3.83		2.62		472		323
Combined Mining
Measured		14,746		1.22		1.84		579		872
Indicated		79,131		1.11		2.26		2,839		5,743
Measured and Indicated		93,877		1.13		2.19		3,418		6,615
Inferred		108,422		0.91		2.32		3,175		8,104

Mineral resources are reported in relation to an elevation determined from optimized pit shells. Mineral resources are not mineral reserves and do not have demonstrated economic viability. All figures are rounded to reflect the relative accuracy of the estimate. All composites have been capped where appropriate.

Open pit mineral resources are reported at a cut-off grade of 0.35 g/t gold and underground mineral resources are reported at a cut-off grade of 2.50 g/t gold. Cut-off grades are based on a price of USD $1,025 per ounce of gold and gold recoveries of 88% and 90% for open pit and underground resources, without considering revenues from other metals.

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In June 2011, SRK prepared a sixth mineral resource statement to incorporate information from an additional 50 core boreholes (26,509 m) drilled on the Project since the last resource model considering data up to February 27, 2011. The sixth consolidated mineral resource statement dated June 29, 2011 is represented in Table 6-6.

Table 6-6: Mineral Resource Statement*, Rainy River Gold Project, Ontario,

SRK Consulting (Canada) Inc., June, 29 2011


	Quantity		Grade				Metal
Category	‘000 t		Au g/t		Ag g/t		Au ‘000 oz.		Ag ‘000 oz.
Open Pit**
Measured		15,660		1.26		1.93		636		973
Indicated		99,927		1.08		2.48		3,481		7,967
Measured and Indicated		115,587		1.11		2.41		4,117		8,940
Inferred (inside pit shell)		16,602		0.94		2.63		504		1,406
Inferred (outside pit shell)		57,211		0.75		2.82		1,380		5,184
Underground**
Measured		100		4.74		2.67		15		9
Indicated		1,775		4.83		3.10		276		177
Measured and Indicated		1,875		4.82		3.08		291		185
Inferred		3,628		3.82		3.84		445		448
Combined Mining
Measured		15,760		1.28		1.94		651		981
Indicated		101,702		1.15		2.49		3,757		8,144
Measured and Indicated		117,462		1.16		2.42		4,407		9,125
Inferred		77,442		0.94		2.83		2,330		7,038

Open pit mineral resources are reported at a cut-off grade of 0.35 g/t gold and underground mineral resources are reported at a cut-off grade of 2.50 g/t gold. Cut-off grades are based on a price of USD $1,100 per ounce of gold and gold recoveries of 88% and 90% for open pit and underground resources, respectively, without considering revenues from other metals.

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In February 2012, SRK prepared a seventh mineral resource statement to incorporate information from an additional 375 core boreholes (181,682 m) drilled on the project since the last resource model considering data up to January 9, 2012. The seventh consolidated mineral resource statement dated February 24, 2012 is represented in Table 6-7.

Table 6-7: Mineral Resource Statement*, Rainy River Gold Project, Ontario,

SRK Consulting (Canada) Inc., February, 24 2012


	Quantity		Grade				Metal
Category	‘000 t		Au g/t		Ag g/t		Au ‘000 oz.		Ag ‘000 oz.
Open Pit**
Measured		23,154		1.29		2.00		960		1,491
Indicated (in pit)		112,778		1.09		2.39		3,963		8,673
Indicated (ex. pit)		11,476		0.81		3.37		298		1,242
Measured & Indicated		147,407		1.10		2.41		5,221		11,406
Inferred (in pit)		22,679		0.93		2.18		675		1,588
Inferred (ex. pit)		64,437		0.67		2.35		1,387		4,871
Underground**
Measured^		89		4.62		2.55		13		7
Indicated^		3,083		4.32		5.00		429		495
Measured and Indicated^		3,172		4.33		4.93		442		502
Inferred^		1,172		4.12		5.82		155		219
Combined Mining
Measured^		23,243		1.30		2.00		973		1,498
Indicated^		127,337		1.14		2.54		4,690		10,410
Measured and Indicated^		150,580		1.17		2.46		5,663		11,908
Inferred^		88,288		0.78		2.35		2,217		6,678

Open pit mineral resources are reported at a cut-off grade of 0.35 g/t gold and underground mineral resources are reported at a cut-off grade of 2.5 g/t gold. Cut-off grades are based on a gold price of USD $1,100 per ounce, and a foreign exchange rate of CAD $1.10 to USD $1.00 and gold recoveries of 88% for open pit and 90% for open pit and underground mineral resources, respectively.

^	Due to a reporting discrepancy, the underground resources reported in the Press Release by Rainy River on February 24, 2012 differ nominally to that reported here.

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7.	GEOLOGICAL SETTING AND MINERALIZATION

7.1

Regional Geology

The Rainy River Gold Project is located within the Late Archean Rainy River Greenstone Belt (“RRGB”) which formed approximately 2.7 billion years ago (“Ga”). The RRGB forms part of the western Wabigoon subprovince, located in the Superior Province of the Canadian Shield. The Wabigoon subprovince is a 900 km long, east-west trending composite volcanic and plutonic terrane comprising distinct eastern and western domains separated by rocks of Mesoarchean age (Percival et al. 2006).

The western Wabigoon domain is predominantly composed of mafic volcanic rocks intruded by tonalite-granodiorite intrusions. The volcanic rocks which were largely deposited between approximately 2.74 and 2.72 Ga, range from tholeiitic to calc-alkaline in composition, and are interpreted to represent oceanic crust and volcanic arcs, respectfully (Percival et al. 2006). These are succeeded by approximately 2.71 to 2.70 Ga volcano-sedimentary sequences and by locally deposited, unconformable, immature clastic sedimentary sequences.

The volcanic rocks have been intruded by a wide variety of plutonic rocks including synvolcanic tonalite-diorite-granodiorite batholiths, younger granodiorite batholiths, sanukitoid monzodiorite intrusions and monzogranite batholiths and plutons. The intrusions were emplaced over a large time span between approximately 2.74 to 2.66 Ga (Percival et al. 2006).

A regional map of the interpreted bedrock geology west of Fort Frances is shown in Figure 7-1. In the region east of Fort Frances, the Wabigoon subprovince is bounded to the south by the late Archean, dextral Seine River–Rainy Lake and Quetico faults. The Quetico Fault splays off the subprovince boundary and strikes west through the western Wabigoon domain just south of the Rainy River Gold Project. The RRGB is bounded to the north by the Sabaskong Batholith and to the east by the Rainy Lake Batholithic Complex and is contiguous with the Kakagi-Rowan Lakes Greenstone Belt to the north.

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The regional metamorphic grade of the Achaean rocks is greenschist to lower-middle amphibolite facies. Locally, adjacent to the intruding batholiths, upper amphibolite mineral assemblages are recognized.

Significant metallic mineral deposits hosted in the western Wabigoon domain include the Cameron Lake gold deposit hosted in the adjacent Kakagi–Rowan Lakes Greenstone Belt, the Hammond Reef gold deposit 190 km to the east of the Rainy River Gold Project, and the Sturgeon Lake Volcanogenic Massive Sulphide (VMS) deposits 250 km to the northeast of the Rainy River Gold Project.

Three (3) phases of the Quaternary Wisconsinan glaciation are recorded in the Rainy River Gold Project (Barnett, 1992). The Archean basement rocks and locally preserved Mesozoic sediments are overlain by till deposited from the Labrador Sector of the Laurentide Ice Sheet. Its provenance area is the Archaen basement of the Canadian Shield to the northeast. In the area of the Rainy River Gold Project, this till has been found to contain highly anomalous concentrations of gold grains, auriferous pyrite and copper-zinc sulphides. As the Labradorean ice sheet retreated, a thick, conductive, geochemically unresponsive glaciolacustrine clay and silt horizon originating from glacial Lake Agassiz was deposited.

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LOGO

Figure 7-1: Regional Bedrock Geology of the Area West of Fort Frances

(modified from Percival and Easton, 2007)

Abbreviations in Figure 7-1 are as follows:

BHS – Black Hawk Stock,

KFLGB – Kakagi-Rowan Lakes Greenstone Belt,

QF – Quetico Fault, RLBC–Rainy Lake Batholithic Complex,

RRGB – Rainy River Greenstone Belt, Rainy River Gold Project,

SRRLF – Seine River–Rainy Lake Fault,

SB – Sabaskong Batholith.

The Keewatin Sector of the Laurentide Ice Sheet then advanced over the area and deposited an argillaceous till of western provenance on top of the clay and silt horizon. The Rainy River Gold Project area was therefore successively covered by the Labradorean and Keewatin ice sheets.

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7.2

Property Geology

The geology of the Rainy River Gold Project is known from regional field mapping of limited rock exposures, extensive drilling by Nuinsco and Rainy River, OGS rotasonic drilling and airborne geophysics.

A bedrock geological interpretation produced by Rainy River for the area surrounding the Rainy River Gold Project is shown in Figure 7-2.

A north-to-south cross section across the Project is provided in Figure 7-3. The Rainy River Gold Project is centred on Richardson Township. To the north of Richardson Township lies the Sabaskong granitoid batholith. The Black Hawk Stock lies to the east of Richardson Township. A package of metasedimentary rock is found south of Richardson Township. Wedged in between these lithologies are a series of tholeiitic mafic and structurally overlying calc-alkalic intermediate to felsic metavolcanic rocks, striking almost east-west and dipping to the south. Intermediate dacitic rocks host most of the Rainy River gold mineralization.

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LOGO

Figure 7-2: Bedrock Geological Interpretation for the Area Surrounding the Rainy River Gold Project (from Rainy River, 2012)

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LOGO

Figure 7-3: A North-South Geological Cross Section Across the Rainy River Deposit (Rainy River, 2012)

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7.2.1

Lithology

Lower Mafic Volcanic Succession

The Rainy River Gold Project area is primarily underlain by tholeiitic metavolcanic rocks of the western Wabigoon subprovince – the Rainy River Greenstone Belt. Geochemically they are high-iron and high-magnesium basalts comprising coarse-grained massive lava flows, massive and pillow flows and flow breccia. Subordinate dacitic tuff and intrusive quartz-feldspar porphyry dikes and sills are commonly interbedded throughout the mafic volcanic rock.

Intermediate-Felsic Porphyritic Intrusive Rock

Swarms of porphyritic intermediate to felsic dikes cut through the Lower Mafic volcanic succession. They range in thickness up to several tens of metres. It has been suggested that these dikes may have been the conduits that fed the overlying intermediate succession hosting the mineralization. They have been variably interpreted and often described as dacitic tuffs due to their similar composition and appearance to units noted within the overlying intermediate succession. Historically, these complex and strongly deformed units have been denoted as the Georgeson/Feeder Porphyries.

Upper Felsic Succession

The Upper Felsic Succession overlies the Intermediate Succession along the southern boundary of Richardson Township. The Upper Felsic Succession is a few hundred metres thick and has been traced for 4 km westwards from the Black Hawk Stock. It has been interpreted as a quartz-phyric rhyolite.

Pinewood Sediment Succession

The Pinewood sedimentary rock package is composed of predominantly clastic intermediate derived wacke and argillite. The sequence conformably overlies the upper diverse mafic volcanic rocks, and the contact is typically marked by a pyritic heavy metal-bearing graphitic horizon. The upper contact of the succession is interbedded with the Upper Felsic Succession.

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Pyritic Sediment Succession

Conformably overlying the Lower Mafic Volcanic Succession are a series of pyrite-bearing siliceous to chloritic wacke, interpreted as derived from intermediate to mafic volcanic sediments. These horizons are increasingly interbedded with homogenous and nondescript to quartz-eye dacite tuff horizons as the upper contact is approached, and these tuff horizons likely represent onset of the lateral equivalent of subsequent intermediate volcanism.

Ultramafic-Mafic Intrusion

Thin zones of ultramafic to mafic intrusions have been noted in drill core. They form dikes or sills intruding the volcanic stratigraphy at different times. Their sulphide content is typically below 2%. The 34 Zone is hosted in a late-stage mafic-ultramafic intrusion, crosscutting the 17 Zone. The main lithological units include dunite, pyroxenite, pyroxene-gabbro and gabbro. The lowermost units contain significant sulphide mineralization enriched in copper, nickel, gold and platinum group metals.

Intermediate Fragmental Volcanic Succession

The Intermediate Succession is complex. Immediately overlying the pyritic sediment horizon in Richardson Township, these volcaniclastic rocks are composed of fine-grained “quartz-eye” dacite and fine-grained ash horizons with subordinate interbedded coarse grained lapilli tuff and localized sedimentary and exhalative horizons. A high proportion of what appear to be coarse volcaniclastic rocks may in fact be massive flows or tuffs overprinted by strong, anastomosing foliation and sericite alteration. Geochemically these intermediate rocks have been interpreted as calc-alkaline dacite with subordinate rhyolite and andesite. Some blocks of tuff breccia have been observed juxtaposed against the Black Hawk Stock which intrudes and notably alters the volcaniclastic rocks to the east. The rocks of the intermediate succession dip 50° to 70° to the south in the Richardson area and are the principal host of the mineralization in the ODM/17, 433, Beaver Pond, Western, and HS Zones.

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Black Hawk Stock

This quartz monzonitic to granodioritic stock consists of two (2) phases and represents a topographic high to the east. The early phase forms the rim of the stock, and is a weakly foliated, notably magnetic, massive to pegmatitic quartz monzonite with minor subordinate granodiorite. The late phase consists of equigranular coarse grained granodiorite, and forms the central core of the stock. Associated magnetic aplitic to pegmatitic dikes compositionally similar to the early phase intrude the surrounding metavolcanic rocks.

Massive Lava Flows

Immediately overlying the intermediate fragmental volcanic rocks are a series of intermediate to mafic volcanic massive lava flows, ranging from fine-grained porphyritic quartz dacite, to massive magnetite-bearing mafic volcanic rocks, with localized pillowed mafic flows. These units are notably homogenous, and the intermediate volcanic units often show a diagnostic deformed sericitic net-textured compression fracture pattern. Upper and lower contacts display a centimetre scale shear fabric at the margins.

Upper Diverse Mafic Volcanics

The upper diverse mafic volcanic succession is composed of a series of mafic tuffs, massive to glomeroporphyritic mafic flows, localized pillowed flows, interflow sediment and hyaloclastite, and minor subordinate intermediate volcanic tuffs. The rocks of the upper diverse mafic volcanics are the principal host of the CAP Zone mineralization.

Proterozoic Diabase Dike

A northwest-striking, steeply dipping diabase dike cross-cuts the ODM/17 Zone and extends across the entire Project area.

7.2.2

Structural Geology

At a regional scale, the strongest and earliest deformation event produced a well-defined penetrative fabric. This foliation is approximately parallel to the trend of the metavolcanic rocks

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which strike at approximately 120° and dip 50° to 70° to the south. This foliation is deformed locally by the intrusion of the Black Hawk Stock. Interpretation of aeromagnetic data (Siddorn, 2007) suggests that a set of tight F1 isoclinal folds with west northwest-trending axial traces cross the area and were then deformed by less intense, north northeast-trending open F2 folds.

Major faults, such as the east-striking dextral Quetico Fault, cross just to the south of the Rainy River Gold Project. West northwest-striking splays off this fault have been interpreted from aeromagnetic data to extend into the Richardson Township area (Figure 7-4).

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LOGO

Figure 7-4: Regional Structural Trends on the Rainy River Gold Project,

Interpreted from Aeromagnetics (Siddorn, 2007)

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Within the Rainy River Gold Project, all rock types examined within the main sulphide mineralized zones except late diabase dikes have a well-developed, moderately south-dipping, penetrative foliation and a moderately southwest-plunging stretching lineation (Figure 7-5 and Figure 7-6).

LOGO

Figure 7-5: Structural Fabrics Affecting Rock Types

A.	Strong, penetrative foliation in sericite-quartz-altered, quartz-phyric rocks. Photo is rotated to approximate dip of borehole (NR09399, 672.2 m).

B.	Southwest raking stretching lineation (NR09402, 245.9 m).

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LOGO

Figure 7-6: Summary of Structural Data Collected by SRK in 2011

Raw (A) and contoured (B) stereonets of 468 poles to foliation measured by Optical Televiewer (OTV) in boreholes NR0501, NR0502, NR0503 NR0504, NR0673, NR0689A, NR06106, NR06107, NR06111, NR06117, NR06134, NR07218, NR09361, NR09367, NR09368, NR09428 and NR09445. Contoured poles to foliation (C) and lineation (D) are measured from oriented drill core in borehole NR0504.

Most phenocrysts have a flattened, symmetrical shape suggesting much of the deformation resulted in a strong flattening and elongation of the host rocks. Kinematic indicators such as asymmetric pyrite grains are also present in the ODM/17 Zone. These suggest that a component of south-over-north reverse-sinistral deformation also affected the rocks (Figure 7-7), which is consistent with a north northwest-south southeast directed compressive deformation event.

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Minor, moderately north dipping, approximately west-striking quartz-calcite-filled brittle-ductile faults have been identified in several boreholes. These faults commonly cause dragging of foliation indicating north over south reverse kinematics. Their orientation is compatible with formation in the overall north northeast-south southeast directed compression event (and may represent a later stage in the protracted deformation history).

LOGO

Figure 7-7: Pressure Shadows Around Rigid Objects in Dacitic Rock from the ODM/17 Zone

(SRK, 2011). Photographs of core from core boreholes, as indicated.

Left - Rock fragment with symmetric pressure shadows.

Right - Asymmetric pressure shadows around pyrite grains, with shear sense, as indicated.

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An approximately 2 to 12-metre-wide High Strain (HS) Zone striking approximately 100° and dipping 55° occurs at the base of the ODM/17 Zone, (Figure 7-8). This zone is characterized by:

•

A penetrative foliation accompanied by strong sericite-chlorite alteration banding and intermittent pervasive silicification;

•

Steepening of dip of foliation surfaces to 75° south;

•

Intermittent kink banding;

•

Generally low gold grades (less than 1.0 g/t gold);

•

Presence of brittle fractures exploiting (commonly kink banded) spaced foliation planes; and

•

Common (deformed) quartz-carbonate-chlorite shear veins near the base of the HS Zone.

LOGO

Figure 7-8: High Strain Zone at the Base of the ODM/17 Zone (SRK, 2011)

A.	Drill core photograph showing development of a HS Zone (approximately 9 m wide) characterized by spaced foliation with sericite-chlorite banding, brittle fractures, pervasive silicification and quartz-carbonate-chlorite shear vein at the base of ODM/17 Zone (NR0643, 329-352 m).

B.	Composite photograph highlighting development of the basal ODM/17 HS Zone (NR0676, 271.8-287.3 m).

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Subsidiary HS Zones are slightly oblique to the HS Zone at the base of the ODM/17 Zone bound panels of gold mineralization and are coincident with small variations in the orientation of the main penetrative foliation between panels (Figure 7-9). The geometry of these subsidiary HS Zones and the associated foliation were used to guide modelling of the medium and high-grade domains within the ODM/17 Zone.

LOGO

Figure 7-9: Oblique 3D View of ODM/17 Zone Looking Down Plunge to the Southwest (SRK, 2011)

Leapfrog modelled gold shells = 0.3, 0.5, and 1.0 g/t gold and foliation measured by an Optical Televiewer are displayed as disks. Small variations in foliation are evident in different panels of the ODM/17 Zone and in its footwall. HS Zones observed in drill core commonly separate these panels.

Subvertical to steeply dipping, approximately east-striking, narrow (less than 1 m wide) chlorite-carbonate shear zones are spatially associated with, and overprint HS Zones in the 433 and ODM/17 Zones and cut the mafic volcanic rocks in the CAP Zone (Figure 7-10). These are associated with shallowly dipping extensional veins and are interpreted as a late component of reverse faulting during a progressive, regional north-northeast-directed compressional deformation event. Quartz-carbonate ±pyrite veins occur along these shears and are commonly folded or boudinaged with pyrite in boudin necks. These veins may carry low-grade gold mineralization in

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the CAP, ODM/17 and 433 Zones, and higher grade gold mineralization when localized in preferential structural sites such as boudin necks.

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Figure 7-10: Chlorite Carbonate Shear Zones (SRK, 2011)

A.	Folded pyrite stringer veins and quartz-carbonate-chlorite shear veins in strongly carbonate-chlorite-sericite-quartz altered mafic volcanic rocks (NR10-474, composite).

B.	Outcrop exposure of folded quartz-carbonate-pyrite veins in foliated carbonate-chlorite-quartz altered mafic volcanic rocks in the CAP Zone (425372E, 5409428N).

At the western end of the ODM/17 Zone, brittle-ductile shear zones have been identified in several boreholes (Figure 7-11) striking at 200° and dipping 75° to the northwest. These anastomosing shear zones are characterized by rotation and crenulation of foliation, intense sericite and chlorite alteration and the presence of quartz-carbonate vein fragments. These shear zones are considered to be responsible for separation (and potential offset) between main ODM/17 Zone and the Beaver Pond subdomain.

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LOGO

Figure 7-11: Brittle Ductile Shear Zones Identified in the ODM/17 Zone

A.	Quartz-carbonate fragments in intensely sericite-chlorite altered rocks in shear zone cross-cutting regional foliation (NR09-414, composite).

B.	Shear zone boundary striking approximately 200 degrees in oriented (relative to lineation) drill core (NR08311, 127.8 m).

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Detailed borehole logging (Ayres, 1997) has found evidence for late, broadly north-south brittle faults offsetting the mineralized horizons. This is supported by field observations (Figure 7-12) by SRK, during site visits in 2007 and 2010 (Siddorn, 2007; Hrabi and Vos, 2010) documenting small-displacement faults with predominantly strike-slip movement, and predominantly sinistral offsets.

LOGO

Figure 7-12: Evidence for Strike-Slip Kinematics of Late Brittle Faulting (SRK, 2011)

A.	Subhorizontal striations on calcite-filled, north-striking subvertical strike-slip faults (borehole NR0505, 150.6 m).

B.	Outcrop (426255E/5409525N) showing set of 360° to 020°-striking, subvertical faults offsetting mafic dike predominantly with sinistral separation in this area.

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SRK structural analyses (Siddorn, 2007; Hrabi and Vos, 2010) have noted that the gold mineralization is strongly overprinted by subsequent deformation.

Key observations in core and outcrop include:

•

Auriferous mineralization is aligned along the regional foliation;

•

Fold axes of auriferous quartz veins and sulphide stringers are rotated subparallel to the stretching lineation;

•

Fold axes, boudin necks and stretching lineation are subparallel to the plunge of the gold mineralization;

•

Early sulphide mineralization is deformed by folding (Figure 7-13); and

•

Later quartz-sulphide veins are variably deformed and overlap in time with the main regional deformation.

LOGO

Figure 7-13: Sulphide Mineralization Deformed by Folding in Core

from the Rainy River Gold Project (SRK, 2011)

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This strongly suggests that the current geometry and plunge of the gold mineralization at the Rainy River Gold Project is the result of high strain deforming features associated with gold mineralization and rotating the ore plunge parallel to the stretching direction (Figure 7-14).

LOGO

Figure 7-14: Structural Control of the Plunge of the Gold Mineralization at the Rainy River Gold Project

A.	Fold axis of pyrite stringer vein rotated parallel to mineral lineation (NR09408, 459.5 m).

B.	Boudinaged quartz vein with boudin neck parallel to stretching lineation (NR09360, 766.4 m).

C.	Diagram illustrating rotation of ore plunge in high strain deformation (modified from Robert and Poulsen, 2001).

7.3

Mineralization

Four (4) main styles of mineralization have been identified on the Rainy River Gold Project:

1.	Moderately to strongly deformed, auriferous sulphide and quartz-sulphide stringers and veins in felsic quartz-phyric rocks (ODM/17, Beaver Pond, 433 and HS Zones);

2.	Deformed quartz-ankerite-pyrite shear veins in mafic volcanic rocks (CAP/South Zone);

3.	Deformed sulphide-bearing quartz veinlets in dacitic tuffs/breccias hosting enriched silver grades; and

4.	Copper-nickel-platinum group metals mineralization hosted in a younger mafic-ultramafic intrusion (34 Zone).

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7.3.1

Auriferous Sulphide and Quartz-Sulphide Stringers and Veins in Felsic Quartz-Phyric Rocks

The bulk of the gold mineralization at the Rainy River Gold Project is contained in sulphide and quartz-sulphide stringers and veins hosted by felsic quartz-phyric rocks. Two (2) main zones are recognized (ODM/17 and 433 Zones) with subsidiary zones (HS and New), which are mostly bound by high strain zones.

ODM/17 Zone

Three (3) styles of gold mineralization can be observed in the ODM/17 Zone. Low-grade intervals are characterized by tightly folded pyrite stringer veins and disseminated pyrite in sericite-quartz-chlorite altered host rocks.

Low- to moderate-grade (up to approximately 10 g/t gold) intervals are characterized by tightly folded and foliation parallel pyrite-sphalerite and pyrite stringer veins, commonly associated with stronger silica and weak garnet alteration (Figure 7-15). High-grade gold mineralization in this zone is associated with deformed quartz-pyrite-gold veinlets (Figure 7-16) that overprint other mineralization styles.

The low-grade ODM/17 Zone is modelled over a strike length of approximately 1,600 m, over a vertical distance of approximately 975 m and over a true width of up to 200 m.

The medium-grade Beaver Pond subdomain (Subdomain 114) is located 300 m to the west of the ODM/17 Zone and is included within the larger ODM/17 low-grade domain. Mineralization is similar in style and character to the ODM/17 Zone, including the presence of deformed gold-rich quartz veinlets, although generally the widths of auriferous drill intercepts are narrower.

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LOGO

Figure 7-15: ODM/17 Zone Gold Mineralization (SRK, 2011)

Deformed pyrite-sphalerite veins and stringers parallel to, or obliquely to foliation in quartz-sericite-chlorite altered rocks (Borehole NR0651 at downhole interval, as indicated).

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LOGO

Figure 7-16: ODM/17 High-Grade Gold Mineralization (SRK, 2011)

Deformed quartz-pyrite vein with visible gold emplaced along boudin neck (Borehole NR0651 at 251.1 m; 195.5 g/t gold over 1 m core length interval).

The 433, HS and New Zones

The style of gold mineralization in the 433 Zone is comparable to that observed in the ODM/17 Zone, although some differences are apparent. These include an overall dominance of chlorite alteration (relative to dominant sericite in the ODM/17 Zone) of quartz-phyric host rocks, occurrences of chlorite-pyrite altered heterolithic conglomerates, and the occurrence of chalcopyrite and chlorite with high-grade quartz-pyrite-gold veinlets (Figure 7-17).

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LOGO

Figure 7-17: 433 High-Grade Gold Mineralization (SRK, 2011)

Deformed quartz-pyrite-chalcopyrite-chlorite-gold veins cross-cutting foliation and disseminated pyrite in quartz-sericite altered quartz-phyric rock (Borehole NR07-218 at 305.2 m; 4,159 g/t gold over 1 m core length interval).

The modelled 433 low-grade Zone has a flattened oblate shape that plunges moderately to the southwest. The zone extends over a strike length of approximately 325 m, over a vertical distance of approximately 820 m and a true width of up to 125 m.

Several subsidiary zones of gold mineralization are identified at the Rainy River Gold Project, including the HS and New Zones. The HS and New Zones are located north of and structurally beneath the ODM/17 Zone and slightly above the 433 Zone. The full extent of the HS Zone has not been defined by drilling to-date.

The HS Zone defines a plunging, flattened oblate shape subparallel to the ODM/17 and 433 Zones which hosts discontinuous, irregular low-grade gold mineralization associated with chlorite-pyrite replacement of matrix in flattened, albitized heterolithic pebble conglomerates. The New Zone is more irregular in shape, comprising a number of small zones in the immediate hanging wall of the 433 Zone. The zones have a strike length of approximately 200 and 275 m respectively, and both extend a vertical distance of approximately 700 m.

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The Western Zone

The Western Zone occurs near surface approximately 1 km northwest of the Beaver Pond Zone. It is composed of stockwork of discrete centimetre scale anastomosing, folded to linear quartz and quartz-carbonate veinlets, hosted predominantly by strongly deformed intermediate volcanic fragmental units, analogous to those that host the ODM/17 Zone, but also present in mafic volcanic flows in both the immediate footwall and hanging wall. The stratigraphy hosting the Western Zone shows a much higher degree of deformation than to the east and combined with intense sericitic alteration and foliation is often described as a pervasive shear fabric or approaching mylonitic texture. The veinlets are variably mineralized, with inclusions (in the order of frequency) of pyrite, anemic sphalerite, chalcopyrite, galena, native silver, electrum and native gold.

7.3.2

Deformed Quartz-Ankerite-Pyrite Shear Veins in Mafic Volcanic Rocks

The CAP Zone

The CAP Zone is located approximately 200 m to the south of the ODM/17 Zone and defines a southwest-plunging lens modelled over a strike length of approximately 400 m to a depth of approximately 400 m below surface. Higher-grade gold mineralization is associated with deformed quartz-ankerite-pyrite shear and extensional veins hosted by quartz-ankerite-pyrite altered mafic volcanic rocks (Figure 7-18). Relative to ODM/17 and 433 Zones, the CAP Zone has a higher pyrite-chalcopyrite content.

7.3.3

Silver-Rich Deformed Sulphide-Quartz Veins within Tuffaceous Rocks

The so-called “Footwall Silver Zone” occurs in altered dacitic tuffs/tuff breccias immediately adjacent to the high strain zone at the northern contact of the ODM/17 Zone. The zone plunges to the southwest in similar orientation to the ODM/17 Zone, and is hosted by centimetre scale sulphide bearing quartz veinlets, typically appearing as millimetre scale fracture filling to dendritic native silver inclusions. Associated sulphides within these veinlets in order of frequency are; pyrite, sphalerite, chalcopyrite, galena. Localized spessartine garnets have been noted. The presence of isoclinal folding of the veinlets gives the mineralization a relative timing of pre to syndeformational, and the zone is currently considered to be coeval with the ODM/17 Zone.

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LOGO

Figure 7-18: Higher Grade Gold Mineralization (SRK, 2011)

Borehole NR10-474 from 188.0 to 234.0 m; gold grade shown over 1 m core-length interval.

7.3.4

Nickel-Copper-PGE Mineralization

The 34 Zone

Magmatic nickel copper sulphide mineralization is found in the 34 Zone. It is associated with precious metals (gold, platinum group metals) and occurs within a tubular, late-stage pyroxenite-gabbro intrusion that crosscuts the ODM/17 Zone. The magmatic sulphides vary from massive to net-textured and disseminated. The host pyroxenite-gabbro intrusion is unmetamorphosed, but locally altered into serpentine and talc. The 34 Zone extends at least 350 m along strike. The host intrusion is approximately 100 m thick and plunges at 12° to the west. The limits of the 34 Zone are poorly constrained owing to the discontinuous nature of the mineralization and a lack of drilling data.

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8.	DEPOSIT TYPES

This section was derived directly from the Preliminary Economic Assessment Update of the Rainy River Gold Property, Ontario, Canada by BBA Inc. published on October 12, 2012. As the information remains valid and relevant, an update to this section was not required.

The origin of gold mineralization at the Rainy River Gold Project remains enigmatic because it has been modified by deformation as highlighted in recent studies (Siddorn, 2007; Hrabi and Vos, 2010).

Early exploration work performed by Nuinsco was based on the premise that the gold mineralization encountered at Rainy River was shear-hosted and epigenetic in origin. Subsequent interpretations for the geology and genesis of the Rainy River Gold Project gold mineralization involved an early, volcanogenic-associated caldera model, as proposed by Ayres (1997) and Averill (2008). Volcanic facies reconstruction studies by Wartman (2011) suggests the presence of a lobe-hyaloclastite dacite dome/flow complex fed and locally intruded by synvolcanic dacite hypabyssal intrusions. Despite the strong alteration overprint, primary textures indicate that the volcanic facies in the deposit include coherent dacitic flows and associated syn-volcanic intrusions with autoclastic breccias, hyaloclastites, peperites and syn- to post-depositional resedimented volcaniclastic deposits.

With the collection of additional data from exploration drilling, it was recognized that gold mineralization at the Rainy River Gold Project is associated with strong sodium depletion, potassium enrichment, aluminous alteration, a strong gold-pyrite association, ubiquitous sphalerite, chalcopyrite and manganiferous garnet (spessartine) and a very high ratio of silver to gold. These features suggest a volcanogenic sulphide origin. However, no significant base metal mineralization or stratiform sulphide lenses have been recognized.

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Siddorn (2007) and Hrabi and Vos (2010) suggested that at least two (2) stages of gold mineralization occurred in the Rainy River Gold Project:

•

Early (low- to moderate-grade) gold mineralization associated with the emplacement of sulphide (pyrite-sphalerite-chalcopyrite-galena) stringers and veins and disseminated pyrite in quartz-phyric volcaniclastic rocks and conglomerates; and

•

Late (high-grade) gold mineralization associated with the emplacement of quartz-pyrite-chalcopyrite-gold veins and veinlets.

Both styles of gold mineralization have been progressively overprinted by deformation, whereby auriferous quartz veins post-date the sulphide stringers and veins and were emplaced during active deformation (Figure 8-1).

On this basis, the gold mineralization is interpreted as a hybrid deposit-type consisting of an early gold-rich volcanogenic sulphide mineralization overprinted by shear-hosted mesothermal gold mineralization.

Volcanogenic deposits refer to a large family of mainly copper-zinc (and subsidiary gold, silver and/or lead) deposit types that are typically related to the precipitation of metals from hydrothermal solutions circulating in volcanically active submarine environments. Volcanogenic deposits typically form during periods of active rifting along volcanic arcs, fore-arcs and in extensional back-arc basins. Gold-rich volcanogenic deposits form a sub-type as described by Dubé et al., (2007).

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LOGO

Figure 8-1: Idealized Sketch Showing Relative Timing of Auriferous Features and Illustrating

Protracted Deformation Affecting Initial Gold-rich Volcanogenic Mineralization and

Subsequently Overprinted by Mesothermal Gold Mineralization (SRK, 2011)

Dubé et al., (2007) indicate that their diagnostic features include stratabound to discordant massive sulphide lenses with associated discordant stockwork feeder zones in which gold is the main commodity. The gold-rich volcanogenic deposits are present in both recent seafloor and deformed and metamorphosed submarine volcanic settings. Several of the largest volcanogenic gold deposits are located in Canada, including the Horne, Bousquet 2-Dumagami, LaRonde Penna and Eskay Creek deposits.

Dubé et al. (2007) proposed that there are two (2) genetic models for gold-rich volcanogenic deposits:

•

Conventional syngenetic volcanic-hosted gold-poor volcanogenic mineralization overprinted during regional deformation by gold mineralization; and

•

Syngenetic volcanogenic deposits characterized by an anomalous fluid chemistry (with magmatic input) and/or deposition within a shallow-water to subaerial volcanic setting

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equivalent to epithermal conditions, in which boiling may have had a major impact on the fluid chemistry.

The deformation and metamorphism that commonly overprint the sulphide mineralization in ancient terranes have obscured the original relationships and led to considerable debate about the syntectonic versus synvolcanic origin of gold-rich volcanogenic deposits, as is also the case at the Rainy River Gold Project.

Gold-rich volcanogenic deposits are characterized by zoned alteration profiles with an outer propylitic (chlorite±sericite) zone surrounding a sericite-rich core representing the “feeder zone” (Figure 8-2). Wartman (2011) and Sparkes & Wartman (2012) have proposed a similar genetic model to the classic model by Hannington et al., 1999, placing the various Rainy River deposits into better context with respect to their stratigraphic position (Figure 8-3).

Although no lenses of massive sulphides have been identified at the Rainy River Gold Project, early pyrite-sphalerite-chalcopyrite-galena stringers and veins may have formed as a stockwork system in the feeder zone (i.e., fault system) as part of a gold-rich volcanogenic deposit.

In addition to the gold mineralization discussed above, the Rainy River Gold Project also contains nickel, copper and platinum group metals sulphide mineralization associated with a differentiated ultramafic-mafic intrusion. That magmatic-hydrothermal mineralization occurs within the main auriferous zones and crosscuts the volcanogenic sulphide mineralization and the later mesothermal gold mineralization associated with the regional deformation.

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LOGO

Figure 8-2: Schematic Geological Setting and Hydrothermal Alteration Associated with Gold-rich

Volcanogenic Hydrothermal Systems (after Hannington et al., 1999)

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LOGO

Figure 8-3: Schematic Section and Hydrothermal Alteration Associated with the

Rainy River Gold Project (Sparkes & Wartman, 2012)

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9.	EXPLORATION

9.1

Period 1967-1989

A summary of the historical exploration work from 1967 to 1989 is discussed in Section 6.1.1.

9.2

Nuinsco Exploration Work (1990-2004)

A summary of the exploration work undertaken by Nuinsco on the Rainy River Gold Project property during 1990 to 2004 is discussed in Section 6.1.2 (refer to Appendix C).

9.3

Rainy River Exploration Work (2005-2012)

In June 2005, Rainy River completed the acquisition of a 100% interest in the Project from Nuinsco.

In the same year, Rainy River relogged key sections of the historical core drilled on the Project property and then input all of the data into a GIS database. Rainy River subsequently drilled in excess of 100 reverse circulation holes in four (4) phases to better define the gold-in-till and gold-in-bedrock anomalies.

Between 2005 and 2007, 218 core boreholes (119,729 m) were drilled. This drilling resulted in the discovery of both the ODM and CAP gold mineralized Zones. The ODM is correlated as being the western extension of the 17 Zone. The CAP is a mineralized zone which is located higher up in the stratigraphy.

In April 2008 CCIC were commissioned to evaluate the mineral resources and prepare a technical report for the Rainy River Gold Project. In 2008, Rainy River drilled an additional 112 core boreholes (59,719 m) and performed a fifth phase of reverse circulation drilling, peripheral to the known gold zones.

In 2009, SRK prepared a mineral resource statement incorporating information from an additional 112 core boreholes (59,719 m) drilled during 2008.

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In early 2010, SRK prepared a revised mineral resource statement to incorporate information from 128 core boreholes (72,164 m) drilled on the Project during 2009.

In early 2011, SRK updated the mineral resource statement to incorporate information from 163 core boreholes (84,648 m) drilled on the project during 2010. In addition, Rainy River Resources discovered the western area mineralization, and continued to delineate the new zone throughout the year. The western area is located approximately 800 m to the northwest of the ODM/Beaver Pond mineralization.

In May 2011, SRK was contracted to prepare a mineral resource statement with an additional 50 core boreholes (26,509 m) completed on the project between December 2010 and February 2011. This mineral resource statement represents the sixth resource evaluation prepared for the Rainy River Gold Project.

Three (3) new significant discoveries were made in 2011. The first was a deep higher-grade gold mineralized ‘shoot’ with a significant silver component, known as the 17 East Extension and represents a potential underground component of the 17 East Zone.

In late 2011, a new high-grade silver zone was discovered in the footwall of the ODM. This zone consists of disseminated to stringer galena mineralization and visible silver. The zone does not host any significant gold (Figure 9-1).

The third discovery in 2011 was a high-grade nickel, copper, PGE discovery, located approximately 1.2 km south of the ODM. This represents the first new magmatic sulphide discovery since the 34 Zone was discovered in 1994. Follow-up drilling and ground exploration is ongoing.

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LOGO

Figure 9-1: Section Showing the Footwall Silver Zone Below the previous PEA Starter Pit

(December 23, 2011) (from Rainy River)

In February 2012, SRK prepared and revised a mineral resource statement with an additional 375 core boreholes (181,682 m) considering data up to January 9, 2012. This mineral resource statement represents the seventh resource evaluation prepared for the Rainy River Gold Project.

In August 2012, exploration drilling discovered a significant new gold and silver zone approximately 1 km east of the proposed open pit boundary of the Rainy River Gold Project. Drill hole NR121258 intersected 2.2 g/t Au and 38.5 g/t Ag over 18.5 m, including 6.0 g/t Au and 83.9 g/t Ag over 3.0 m at a vertical depth of 210 m. This new zone, termed the Intrepid Zone, clearly demonstrates the potential for new discoveries along strike of known mineralization. The new zone was discovered after Mobile Metal Ion (“MMI”) geochemistry identified areas of anomalous gold over the prominent

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magnetic low trend that hosts the majority of the Rainy River Gold Project deposits. The new zone contains disseminated and fracture-related mineralization including 2-3% pyrite and variable amounts of sphalerite, galena and chalcopyrite. Both electrum and visible gold have also been identified in core. A total of 102 holes have now penetrated the Intrepid Zone over a 410 m strike length, and have traced the mineralization down dip for 450 m. Wide zones of near-surface gold mineralization continue to contribute mineralization that have the potential to be exploitable via open pit mining, complemented by higher grade intersections that have the potential to be exploitable via underground mining early in the mine life. Four diamond drill rigs will continue to focus on expanding the thickest portions of Intrepid in the plunge direction, and on better defining the zone in preparation for the release of a National Instrument 43-101 compliant resource in the third quarter (Q3) of 2013. The zone provides a target for further drill delineation and a potential future mining opportunity for Rainy River.

In October 2012, SRK prepared and revised a mineral resource statement with an additional 237 holes (95,760 m) considering data up to July 2012, and excluding the Intrepid discovery. This mineral resource statement represents the eighth resource evaluation prepared for the Rainy River Gold Project and is reported in this Technical Report.

To-date, the Rainy River Gold Project hosts several significant gold mineralized zones stretching over a 3.5 km strike length comprising (from west to east), Western Area, Beaver Pond, ODM, 17, 17 East, 17 East Extension, 280 Zone and the newly discovered Intrepid Zone. The HS, New and 433 Zones are located north of the ODM, while the CAP Zone is to the south.

The exploration work undertaken by Rainy River between 2005 and 2012 is summarized in Table 9-1.

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Table 9-1: Summary of Exploration Work by Rainy River on the

Rainy River Gold Project between 2005 and 2012


Activity	Date	Performed by
2005
Re-Log 13 Diamond Drill Holes	Feb 2005	L.D. Ayres
Summary of Structural Observations	Feb 2005	G. Zhang
Re-Log 8 Nuinsco Diamond Drill Holes	April 2005	L.D. Ayres
53 RC Holes - Phase 1 Drilling	April 2005	Overburden Drilling Management
Technical Report	April 2005	Clark Exploration Consulting
Re-Log 24 Nuinsco Diamond Drill Holes	June 2005	L.D. Ayres
31 RC Holes - Phase 2 Drilling	Aug 2005	Overburden Drilling Management
Petrography and Mineralogy	Aug 2005	R.P. Taylor
DD Drill 17 Holes (3,800 m) BP and W Zones	Sept-Dec 2005	C.J. Baker, N. Pettigrew
Structure and Geology of Caldera Model	Oct 2005	L.D. Ayres
Structure and Geology of Richardson Twp	Oct 2005	H. Paulsen
22 RC Holes - Phase 2 Drilling	Dec 2005	Overburden Drilling Management
2006
Report of Re-Logging of Nuinsco DD Core	Jan 2006	L.D. Ayres
DD Drill 121 Holes (55,219 m)	Jan-Dec 2006	W.Rayner, B. Nelson, A. Tims
Vtem Airborne Geophys. Survey	March 2006	Geotech Limited
U-Pb Zircon Age Dating	June 2006	Geospec Consultants Limited
Petrographic and Mineralogical Report	June 2006	E. Schandl
Structure and Geology Review	July 2006	K. H. Paulsen
U-Pb Zircon Age Dating	Nov 2006	Geospec Consultants Limited
3D Borehole Pulse EM Survey	Nov-Dec 2006	Crone Geophysics and Exploration
2007
DD 179 Holes in 17/ODM, 433 and BP Zones	Jan-Dec 2007	B. Nelson, A. Tims, R. Greenwood
Site Visit by Geologist	Jan 2007	Panterra Geoservices Inc., D. Rhys
IP Survey of 9 Holes, 3D Cond. Inv. Models	May, 2007	JVX Limited
Line Cutting	Nov-Dec 2007	Archer Exploration Inc.
Ground Gravity and EM Survey	Dec 2007	Abitibi Geophysics
2008
DD 102 Holes (53,733 m) 17/ODM and 433 Zone	Jan-Dec 2008	W. Rayner, B. Nelson, A. Tims, C. Hercun
Titan 24 Survey	Jan 2008	Quantec Geoscience
Airborne Magnetic Gradiometer Survey	Feb 2008	Fugro Airborne Surveys, Corp.
47 RC Holes - Phase 5 Drilling	Apr-Jun 2008	Overburden Drilling Management
Regional Geophysical Interpretation	May 2008	J. Siddorn - SRK
Socio-Economic Scoping Study Draft Report	Nov 2008	Klohn, Crippen and Berger Ltd.
Preliminary Pit Slope Design and Waste Management Assessment	Nov 2008	Klohn, Crippen and Berger Ltd.

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Activity	Date	Performed by
2009
DD 139 Holes (69,753 m) in 17/ODM, 433, CAP, South and BP Zones	Jan-Dec 2009	W. Rayner, B. Nelson, A. Tims, C. Hercun, A. Shute
Phase 5, 28-Hole RC Program	Jun-Aug 2009	Overburden Drilling Management
Preliminary Metallurgical Testing	Sep-Nov 2010	SGS Canada Inc.
Preliminary Pit Slope Stability Assessment	Nov 2008- March 2009	Klohn Crippen Berger Ltd.
Preliminary Pit Slope Design and Waste Management Assessment	Nov 2008- Feb 2009	Klohn Crippen Berger Ltd.
Acid Leach Test	Jun-Aug 2009	Klohn Crippen Berger Ltd.
Socio-Environmental Baseline Assessment.	May-Oct 2009	Klohn Crippen Berger Ltd.
Surficial Drainage Project	Apr-Sept 2009	K. Smart Associates Limited
Lidar Survey	Sept 2009	Lidar Services International
New Office Building	Jul-Oct 2009	W. Rayner, K. Schram, B. Burnell, R. Burnell
Third Mineral Resource Statement	Apr 2009	SRK Consulting (Canada) Inc.
Age Dating of Lithologies	Jan-Feb 2009	University of Toronto Geochronology Lab
2010
DD 196 holes (88,61 m) in 17/ODM, 433, CAP, South and BP Zones	Jan-Dec 2010	W. Rayner, B. Nelson, A. Tims, C. Hercun, A. Shute, J. Pattison, H. Buck, K. Pedersen
Deepened 48 Holes (17,718 m) in the 17/ODM Zones	May-Oct 2010	W. Rayner, B. Nelson, A. Tims, C. Hercun, A. Shute, J. Pattison, H. Buck, J. Wartman
DD 4 Geotechnical Drill Holes (1,405 m)	Mar-May 2010	Klohn Crippen Berger Ltd.
NI 43-101 Mineral Resource Estimate	Feb 2010	SRK Consulting (Canada) Inc.
Infill Sampling of Historical Drill Holes - 9504 Samples (13,660 m)	Jul-Aug 2010	A. Tims, C. Hercun, A. Shute, H. Buck, J. Pattison
Phase 6, 37-hole RC Program (1,066 m)	Jan-Feb 2010	Overburden Drilling Management
Review of Pit Slope Design	Mar-May 2010	SRK Consulting (Canada) Inc.
Phase 7, 34-Hole RC Program (762 m)	Sept-Nov 2010	Overburden Drilling Management
New Core Logging Facility	Jul-Dec 2010	C. Hercun, True-line Construction
M.Sc Thesis on Richardson Deposit	Jun-Sept 2010	J. Wartman - University of Minnesota
Pre-Feasibility Open Pit Slope Design	Jun-Dec 2010	Klohn Crippen Berger
Metallurgical Testwork	Jan-Dec 2010	SGS Canada Inc.
Environmental Baseline Studies	Jan-Dec 2010	Klohn Crippen Berger
Memorandum of Understanding with Fort Francis Chiefs Secretariat	May 2010	Rainy River Resources Ltd.
Structural Study	May-Sept 2010	SRK Consulting (Canada) Inc.
Line Cutting Geophysical Grid 33 km	Nov-Dec 2010	Archer Exploration Inc.
Titan Survey 33 km	Dec 2010	Quantec Geoscience
Application for Advanced Exploration Permit	Dec 2010	G. Macdonald, K. Stanfield

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Activity	Date	Performed by
2011
DD 157 holes (77,222 m) in 17/ODM, 17 East, 433, The Gap, South, Beaver Pond, District, Northern, and HS Zones	Jan-Jun 2011	A. Tims, C. Hercun, A. Shute, H. Buck, J. Pattison, K. Pedersen, A. Philippe, R. Montufar, M. Rousseau.
28 Overburden Holes (618 M) Over First 600 m of Proposed Ramp.	Jan-Feb 2011	A. Tims, C. Hercon, A. Shute,H. Buck, J. Pattison, K. Pederson, A. Philippe
88 km High-Sensitivity Potassium Magnetometer Ground Survey	Jan-Feb 2011	RDF Consulting
Fourth Mineral Resource Statement	Feb 2011	SRK Consulting (Canada) Inc.
Fifth Mineral Resource Statement	April 2011	SRK Consulting (Canada) Inc.
Fugro AEM Survey	May 2011	Fugro Airborne Surveys Corp.
Environmental Baseline Gap Analysis	Apr-May 2011	AMEC Earth and Environmental
First Quarter QA/QC Report	Apr-June 2011	Analytical Solutions Ltd.
Sixth Mineral Resource Statement	June 2011	SRK Consulting (Canada) Inc.
Report on Ground Gravity Surveys	Sept-Oct 2011	Eastern Geophysics, Gerard Lambert
Report on Borehole Surveys	Sept-Oct 2011	Eastern Geophysics, Gerard Lambert
2012
2011 Rainy River Gold QC Report	Jan 2012	Analytical Solutions Ltd.
Seventh Mineral Resource Statement	Feb 2012	SRK Consulting (Canada) Inc.
DD 237 Holes (95,760 m)	Jan-Jul 2012	Rainy River Resources Ltd.
Eighth Mineral Resource Statement – Reported herein.	Oct 2012	SRK Consulting (Canada) Inc.

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10.

DRILLING

10.1

Drilling from 2004 to 2011

10.1.1

Introduction

Nuinsco and Rainy River Resources drilled a combined total of 1,435 core boreholes (662,849 m) between 1994 and 2012 (refer to Table 10-1). Prior to 1999, Nuinsco also drilled several reverse circulation boreholes to sample basal till and bedrock for exploration targeting. Reverse circulation drilling data was not used for resource estimation. The mineral resources reported herein consider the drilling data available to July 10, 2012.

The distribution of the core drilling is shown in Figure 10-2 and is summarized in Table 10-1 and Figure 10-1. The Nuinsco data contributes about 9% of the total drilling information (measured by drill metres). Rainy River Resources’ drilling data represents 91% of the drilling data.

Table 10-1: Core Drilling Completed on the Rainy River Gold Project (1994-2012)


Company	Period		Type	No. of Holes		Total Length (m)
Nuinsco		1994 - 2004	Core		199		49,351
Rainy River Resources		2005 - 2007	Core		218		119,729
		2008	Core		100		53,123
		2009	Core		128		72,164
		2010	Core		165		84,134
		2011 - 2012	Core		388		188,588
		2012	Core		237		95,760

Total		1994 - 2012	Core		1,435		662,849

Note: Table reflects all exploration drilling and will differ from the total meterage incorporated into resource calculations.

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LOGO

Figure 10-1: Core Drilling Data by Period (1994 to 2012)

LOGO

Figure 10-2: Drill Collar Map in Relation to Resource Domains and Conceptual Pit Outline

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10.1.2

Drilling Procedures

Rainy River drill programs are designed and conducted by an experienced exploration team under the supervision of a Project Manager and a Vice President, Exploration. The drill procedures used by Nuinsco are not well documented. SRK cannot comment on the procedures used by Nuinsco.

Nuinsco drilling focused on the ODM/17 and 433 Zones, whereas Rainy River drilling focused on infill and step-out drilling in these two (2) zones and on the investigation of new zones (such as the CAP and HS Zones). The discussion below focuses on the drilling procedures adopted by Rainy River.

Most of the recent drilling has been completed by Bradley Bros. Ltd. All drilling used NQ core tools from surface collars. Active drilling was taking place at the time of SRK’s site visits.

A hand-held GPS is used initially to locate and prepare drilling pads in the field. After each hole is complete, a Differential Global Positioning System (“DGPS”) is used to survey the casing with typical accuracy in the tens of centimetres range. The DGPS accuracy is validated on a known control station.

The path of the core borehole is surveyed using a Reflex EZ-SHOT™ instrument, which is an electronic solid-state, single-shot drill hole survey tool, at downhole intervals of 50 m. Their path typically flattens up with depth and wanders off section on the deeper boreholes. Borehole deviation is regarded as a critical issue, as the average borehole length is approximately 400 m. In SRK’s opinion, the survey method used by Rainy River is appropriate. In later 2011, Rainy River started using Tech Directional Drilling to help steer the deeper holes for better control and targeting of the zones. This program was successful and implemented for almost a year.

Rainy River uses a well-designed procedure for logging the drill core and the subsequent integration of this information into the exploration database. Core logging is recorded directly onto laptop computers equipped with DHLogger logging software which ensures that all relevant information is consistently captured and transferred to the main database. Descriptive geological information is recorded with the appropriate validation procedures in place. After validation, logging

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information is transferred directly from the DHLogger database into Gemcom Project software for 3D visualization, interpretation and modeling.

Standardized logging procedures include the collection of lithological, structural, mineralization and alteration features. Magnetic susceptibility readings are recorded every 3 m. Core recovery is reported to be excellent, but was not measured until 2008, when procedures were upgraded to include core recovery and enhanced geotechnical measurements. Geotechnical parameters such as Rock Quality Designation (“RQD”), joint/fracture analyses, material type and rock strength are measured on selected orientated core. Pre-2008 structural and geotechnical logging was usually not based on orientated core. More recently, Rainy River submitted several samples per borehole for specific gravity analyses.

Core is not routinely photographed, although significant intersections and features are photographically recorded. Diamond core is archived in secure, high quality storage facilities on the main Rainy River Gold Project site.

10.2

Drilling Pattern and Density

Core boreholes are mostly angled boreholes drilled on northerly directed azimuths, predominantly at a dip angle of 50 to 55 degrees. The main zones of gold mineralization have been drilled on at least 60 m x 60 m centres (Figure 10-2). The ODM/17 Zone was investigated on a more detailed 30 m x 30 m grid pattern.

10.3

SRK Comments

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11.

SAMPLE PREPARATION, ANALYSES, AND SECURITY

There are no records describing the sampling method and approach used by Nuinsco during their 1994 to 2004 drilling program. When SRK visited the Rainy River Gold Project property, drilling and sampling was active. The following information regarding Rainy River sampling procedures is derived from direct observations as well as from discussions with Rainy River personnel.

Standardized core sampling protocols are used by Rainy River. Initially, Rainy River selectively sampled each core borehole based on visible observation of mineralization and alteration. Core was marked for sampling at regular 1.5 m intervals and half core was cut and sampled. Since then, sampling is performed for the entire length of each hole. The maximum, and most common, sample interval is still 1.5 m. Shorter samples are collected to demarcate geological domains, however, a minimum sample interval of 1 m is maintained.

The sampling interval is the last item marked on the core and recorded in the log. A qualified geologist marks out sample intervals with a red grease pencil and places two (2) sample tags at the beginning of each sample interval. A third copy of the sample tag remains in the sample booklet along with “from” and “to” information recorded by the geologist. These tags are kept in the main office and filed with each individual hole.

The core boxes with samples marked are then prepared for cutting. Once a sample is cut, one half of the core is rinsed and placed into a sample bag and the second half is returned to the core box. One of the sample tags is placed in the sample bag, while the other remains in the core box for reference. The sample bags are stapled closed by the core cutting technician and also individually marked with each sample number. Five (5) sample bags are normally placed into a labelled rice bag and stored in a secured area prior to dispatch to the assaying laboratory. Each hole is separated by placing the rice bags on separate wooden pallets, never combining holes on one pallet.

Sample shipments are typically coordinated two (2) days per week, to make sure the shipment is never left overnight or over weekends at the shipping yard. A photocopy of the sample submission form is placed inside the first rice bag of each hole. The rice bags are transported directly to

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Gardewine North Shipping, in Fort Frances. A typical dispatch contains approximately 400-600 samples. Rice bags that required storing overnight are securely stored inside a designated building.

Following completion of core cutting and sample packing, the core boxes containing the remaining half core are stored outdoors, on sheltered racks. Unsampled intervals in the Nuinsco boreholes were subsequently sampled by Rainy River and incorporated into the borehole database.

In the opinion of SRK, the sampling methodology and procedures used by Rainy River are appropriate. The core samples were collected by competent personnel using procedures meeting with generally accepted industry best practices. SRK concludes that the samples are representative of the source materials and there is no evidence of bias.

11.1

Sample Preparation and Analyses

11.1.1

Nuinsco Samples

There are no records describing the sampling preparation, analyses and security approach adopted by Nuinsco in their 1994 to 2004 programs. It is not known if analytical quality control measures were implemented.

11.1.2

Rainy River Samples

Between 2005 and 2012, Rainy River collected core samples from 1,236 core boreholes (613,498 m). The following sections describe the sample preparation, analyses and security during this period. Core samples were submitted to three (3) separate primary laboratories for preparation and assaying:

•

2005-2006 and since February 2011: ALS Minerals, North Vancouver, British Columbia;

•

2006-2011: Accurassay Laboratory, Thunder Bay, Ontario; and

•

2009: Activation Laboratories, Thunder Bay, Ontario.

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From early 2005 to late 2006 Rainy River used the ALS Minerals Laboratories (“ALS”) in Thunder Bay, Ontario for sample preparation and ALS Minerals Laboratories (“ALS”) North Vancouver, British Columbia for analyses. The management system of the ALS Group Laboratories is accredited ISO 9001:2000 by QMI Management Systems Registration. The North Vancouver Laboratory is accredited ISO/IEC 17025:2005 for certain testing procedures including those used to assay samples submitted by Rainy River. ALS Laboratories also participated in international proficiency tests such as those managed by CANMET and Geostats Pty Ltd.

Between late 2006 and early 2011, Rainy River utilized, almost exclusively, the Accurassay Laboratory (“Accurassay”) facility in Thunder Bay as a primary laboratory. On February 27, 2002, the Standards Council of Canada accredited Accurassay Laboratories for certain testing procedures under ISO/IEC Guideline 17025. The scope of accreditation includes fire assay analyses with atomic absorption finish for gold, platinum and palladium, as well as aqua regia digestion with atomic absorption finish for copper, nickel and cobalt.

Rainy River used the ALS Minerals Laboratory in North Vancouver, British Columbia as an umpire laboratory to monitor the reliability of assaying results delivered by Accurassay during 2010.

In late 2009, Rainy River also used Activation Laboratories (“Actlabs”) in Thunder Bay, Ontario to accelerate the delivery of a small amount of sample batches prior to the resource estimate update of March 2010. Actlabs is also accredited ISO/IEC Guideline 17025:2005 by the Standards Council of Canada for certain testing procedures including gold and silver assaying using a fire assay procedure.

In February 2011, Rainy River reverted to ALS as the primary laboratory for the Project.

The general sample preparation and analyses procedures used by ALS during 2005 – 2006 are described in a previous Technical Report (CCIC, 2008). A description of the sample preparation and analyses procedures adopted by Accurassay (2006 – 2011) and ALS Minerals (2011 – 2012) is provided herein.

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Accurassay Laboratory (2006 – 2011)

Sample Preparation

Samples are first entered into a local information management system (“LIMS”). The protocol for sample preparation at Accurassay involves drying, crushing, splitting, pulverizing and matting:

•

Drying: Prior to the preparation of drill core, the samples are placed in a drying oven, if necessary (approximately 50°C), until dry;

•

Crushing: The entire sample is crushed using a TM Engineering Rhino Jaw crusher to below 10 mesh;

•

Splitting: Approximately 500 g subsamples are split off using a Jones Riffle Splitter;

•

Pulverizing: Samples are pulverized using a TM Engineering ring and puck pulverizer with 500 g bowls to 90% below 150 mesh (105 microns). The bowls are cleaned with silica sand between each sample; and

•

Matting: Pulverized samples are matted to ensure homogeneity.

The homogeneous sample is then sent to the fire assay laboratory or the wet chemistry laboratory depending on the analysis required.

Precious Metal Analyses

Precious metal analyses (gold, platinum, palladium and/or rhodium) require that the sample is mixed with a lead-based flux and fused for one (1) hour and 15 minutes. Each sample has a silver solution added to it prior to fusion, which allows each sample to produce a precious metal bead after cupellation. The fusing process produces lead buttons that contain all of the precious metals from the sample as well as the silver that was added.

The button is then placed in a cupelling furnace where all of the lead is absorbed by the cupel and a silver bead, which contains any gold, platinum and palladium, is left in the cupel. The cupel is removed from the furnace and allowed to cool. Once the cupel has cooled sufficiently, the silver bead is placed in an appropriately labelled test tube and digested using aqua regia. The samples are bulked up with 1.0 mL of distilled de-ionized water and 1.0 mL of 1% digested lanthanum solution. The samples are allowed to cool and are mixed to ensure proper homogeneity of the solution.

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Once the samples have settled, they are analyzed for gold, platinum and palladium using atomic absorption spectroscopy. The atomic absorption spectroscopy unit is calibrated for each element using the appropriate ISO 9002 certified standards in an air-acetylene flame. The results for the atomic absorption are checked by the technician and then forwarded to data entry by means of electronic transfer and a certificate is produced. The laboratory manager checks the data, validates the certificates and issues the results in the format requested by Rainy River.

Base Metal Analyses

Base metal samples (copper, nickel, cobalt, lead, zinc, and silver) are weighed for a geochemical analysis and digested using aqua regia. The samples are bulked to a final volume and mixed. Once the samples have settled they are analyzed for copper, nickel and cobalt using atomic absorption spectroscopy. The atomic absorption spectroscopy unit is calibrated for each element using the appropriate ISO 9002 certified standards in an air-acetylene flame. The results for the atomic absorption are checked by the technician and then forwarded to data entry by means of electronic transfer and a certificate is produced. The laboratory manager checks the data and validates the certificates and issues the results in the format requested by Rainy River.

Analytical Quality Control Measures

Accurassay employs an internal quality control system that tracks certified reference materials and in-house quality assurance standards. Accurassay uses a combination of reference materials, including reference materials purchased from CANMET, standards created in-house by Accurassay and tested by round robin with laboratories across Canada, and ISO certified calibration standards purchased from suppliers.

Should any of the standards fall outside the warning limits (± two (2) standard deviations); re-assays will be performed on 10% of the samples analyzed in the same batch and the re-assay values are compared with the original values. If the values from the re-assays match original assays the data is certified, if they do not match, the entire batch is re-assayed. Should any of the standards fall outside the control limit (± three (3) standard deviations) all assay values are rejected and all of the samples in that batch will be re-assayed.

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ALS Minerals (2011 - 2012)

ALS undertakes precious metal analyses by fire assay with an atomic absorption spectroscopy (“AAS”) finish.

Sample Preparation

The sample is logged in the tracking system, weighed, dried and finely crushed to better than 70% passing a 2 mm (Tyler 9 mesh, US Std. No.10) screen. A split of up to 250 g is taken and pulverized to better than 85% passing a 75 micron (Tyler 200 mesh, US Std. No. 200) screen. This method is appropriate for rock chip or drill samples. ALS sample preparation method codes applied are: LOG-22, DRY-21, CRU-31, SPL-21 and PUL-31.

Precious Metal Analyses

Sample decomposition is by fire assay fusion (ALS method codes FA-FUSO1 and FA-FUSO₂), whereas the analytical method is Atomic Absorption Spectroscopy (ALS method codes Au-AA23 and Au-AA24).

A prepared sample is fused with a mixture of lead oxide, sodium carbonate, borax, silica and other reagents as required, inquarted with 6 mg of gold-free silver and then cupelled to yield a precious metal bead. The bead is digested in 0.5 mL dilute nitric acid in the microwave oven, 0.5 mL concentrated hydrochloric acid is then added and the bead is further digested in the microwave at a lower power setting. The digested solution is cooled, diluted to a total volume of 4 mL with demineralized water, and analyzed by atomic absorption spectroscopy against matrix-matched standards.

Samples grading over 10 g/t gold were analyzed by gravimetric methods (ALS method codes Au-GRA21 and Au-GRA22).

Analyses of Other Metals

ALS also undertakes multi-element analyses by inductively coupled plasma with atomic emission spectroscopy (ICP-AES).

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Sample decomposition is by HF-HNO3-HClO4 acid digestion, HCl leach (ALS method code GEO 4A01), whereas the analytical method is ICP-AES or inductively coupled plasma - mass spectrometry (ICP-MS).

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 analyzed by ICP-AES. Following this analysis, the results are reviewed for high concentrations of bismuth, mercury, molybdenum, silver and tungsten and diluted accordingly. Samples meeting this criterion are then analyzed by ICP-MS. Results are corrected for spectral inter-element interferences.

11.1.3

Metallurgical Testing

Rainy River used the SGS Canada Minerals Services Lakefield Laboratory in Lakefield, Ontario (Lakefield) for metallurgical testing work. The Lakefield Laboratory is accredited ISO/IEC 17025:2005 for certain testing procedures including those used to test and assay samples submitted by Rainy River. The Lakefield Laboratory also participated in international proficiency tests such as those managed by CANMET and Geostats Pty Ltd. The metallurgical testwork completed by Lakefield is discussed in more detail in Section 13.

11.2

Specific Gravity Data

The specific gravity database contains 11,827 measurements completed by Accurassay Laboratory and more recently ALS by pycnometry on pulverized split core samples selected as representative of each modelled geological domain. SRK has composited the specific gravity data to 1.5 m length intervals. Figure 11-1 shows boxplots of the specific gravity composites in each domain.

There are sufficient data to use interpolation, a specific gravity field in the block model for the ODM/17, 433, HS and CAP domains. For all other domains, SRK assigned an average specific gravity value as indicated in Table 11-1.

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LOGO

Figure 11-1: Summary of Specific Gravity Composite Data. Top: All Resource Domains;

Middle: ODM/17 Zone Sub-Domains; and Bottom: Other Domains (600 Domains)

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Table 11-1: Specific Gravity Assigned to Gold Zones


Zone	Domain Code	Specific Gravity
ODM/17	100, 110-115, 120-123	Interpolated
Zone 433	300, 310,3 20	Interpolated
Zone 34	200	2.99
HS Zone	400	Interpolated
CAP Zone	500	Interpolated
Intermediate Zones (600 Series)	601	2.77
	602	2.78
	603	2.77
	604	2.84
	605	2.87
Western Zone	800	2.88
Silver Zone 1	901	2.84
Silver Zone 2	902	2.87
Silver Zone 3	903	2.81
Silver Zone 4	904	2.70

11.3

Quality Assurance and Quality Control Programs

Quality control measures are typically set in place to ensure the reliability and trustworthiness of exploration data. These measures include written field procedures and independent verifications of aspects such as drilling, surveying, sampling and assaying, data management and database integrity. Appropriate documentation of quality control measures and regular analysis of quality control data are important as a safeguard for project data and form the basis for the quality assurance program implemented during exploration.

Analytical control measures typically involve internal and external laboratory control measures implemented to monitor the precision and accuracy of the sampling, preparation and assaying. They are also important to prevent sample mix-up and to monitor the voluntary or inadvertent contamination of samples. Assaying protocols typically involve the use of quality control samples (blank, duplicate samples, certified reference material) to monitor the reliability of assaying results delivered by the Assay Laboratory. Check assaying is normally performed as an additional test of

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the reliability of assaying results. This typically involves re-assaying a set number of sample rejects and pulps at a secondary umpire laboratory.

The review of analytical quality control measures implemented prior to December 2011 is discussed in previous Rainy River technical reports. This Technical Report reviews the analytical quality control measures implemented by Rainy River between December 2011 and June 2012.

Between December 2011 and June 2012, Rainy River relied partly on the internal analytical quality control measures implemented by ALS Minerals. In addition, Rainy River implemented external analytical quality control measures on all sampling consisting of using control samples (blanks, certified reference material and field duplicates) inserted in all sample batches submitted for assaying at a rate of one (1) control sample every 25 samples.

Field blanks consisted of crushed rock material sourced locally from the local Black Hawk Stock and homogenized pulp prepared by Accurassay Laboratories and a second field blank which consisted of a mixture of finely pulverized feldspars and basalt, sourced from Rocklabs Ltd. (Rocklabs), New Zealand, named AuBlank36A and certified as less than 0.002 ppm gold. In May 2011, the field blanks were changed to a medium-hardness coarse marble garden stone sourced from Quali-Grow Garden Products Inc., Canada.

Eight (8) commercial certified reference material control samples were used during this period, sourced from CDN Resource Laboratories Ltd. (“CDN”), Canada.

11.4

SRK Comments

In the opinion of SRK, Rainy River personnel used care in the collection and management of field and assaying exploration data. The analysis of the analytical quality control data is presented in Section 12.

In the opinion of SRK, the sampling preparation, security and analytical procedures used by Rainy River are consistent with generally accepted industry best practices and are therefore adequate.

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12.

DATA VERIFICATION

12.1

Verification of Nuinsco Data

Considering the lack of documented exploration procedures adopted during the Nuinsco exploration program, CCIC (2008) recommended that Rainy River re-sample Nuinsco cores for verification of the assays. According to CCIC (2008), re-sampled core assays compare well with original assay results reported by Nuinsco.

12.2

Verifications by Rainy River

As outlined in Section 11.3, Rainy River relied partly on the internal analytical quality control measures implemented by the accredited ALS Mineral, Accurassay and Actlabs Laboratories but also implemented external analytical quality control measures consisting of inserting control samples into all sample batches submitted for assaying and requesting replicate pulp assays every 12^th sample.

During drilling, experienced Rainy River geologists implement industry standard measures designed to ensure the reliability and trustworthiness of the exploration data. During 2011, Rainy River strengthened the quality control measures partly in response to the recommendations of SRK (2011a). These measures designed by Analytical Solutions include protocols for drill core sampling, sample dispatch, insertion of quality control materials, assessment of incoming data for quality control failures, actions required to remediate quality control failures, regular submission of sample pulps for check assays and collection of core and preparation duplicates. Analytical Solutions also reviewed quality control data up to December 2011 and recommended procedures for ongoing monitoring of analytical results as submitted by the primary laboratory, investigations of dubious results and preparation of regular quarterly quality control reports. SRK considers that the improved procedures adequately address the deficiencies noted previously (e.g., SRK, 2011a) and note that the enhanced procedures have positively impacted the SRK’s analyses of the analytical quality control data.

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12.3

Verifications by SRK

12.3.1

Site Visit

In accordance with the National Instrument 43-101 Guidelines, Glen Cole, P.Geo, visited the Project site on numerous occasions since 2008, most recently from April 30 to May 2, 2013. He was accompanied by various Rainy River staff. At the time of the site visits, active drilling was taking place. The purpose of the site visits was to ascertain the geological setting of the Rainy River Gold Project, witness the extent of exploration work carried out on the Project property, and assess logistical aspects and other constraints relating to conducting exploration work in this area.

All aspects that could materially impact the integrity of the resource estimate (such as core logging, sampling and database management) were reviewed with Rainy River staff. SRK was given full access to all relevant project data. SRK was also able to interview exploration staff to ascertain exploration procedures and protocols.

The location of several borehole collars in the ODM/17, 433 and CAP Zones were verified in the field by SRK. The collars are clearly marked by casings with caps inscribed with the borehole number. No discrepancies were found between the location, numbering or orientation of the holes verified in the field and on plans and the database examined by SRK.

Blair Hrabi, P.Geo, Dr. Ivo Vos, P.Geo, and Simon Craggs from SRK, examined cores from numerous boreholes during various site visits (drilled by Nuinsco and Rainy River) and found the logging information to accurately reflect core intervals. The lithology and gold mineralization contacts checked by SRK match the information reported in the drill logs. Generally, the boundaries of the auriferous zones examined in cores match the boundaries determined from assay results.

12.3.2

Verifications of Analytical Quality Control Data

The analysis of analytical quality control data produced by Rainy River prior to March 2011 was discussed in previous technical reports and is not reproduced here. Rainy River provided SRK with

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external analytical control data in the form of Microsoft Excel spreadsheets containing the assay results for the quality control samples for the period of December 2011 to July 2012.

SRK aggregated the assay results of the external analytical control samples for further analysis. Control samples (field blanks and certified standards) were summarized on time series plots to highlight the performance of the control samples. Paired data (field duplicates and check assays) were analyzed using bias charts, quantile-quantile, and relative precision plots. These plots and charts are presented in Appendix D.

The external analytical quality control data produced by Rainy River from December 2011 to July 2012 are summarized in Table 12-1. The external quality control data produced during 2011 and 2012 represents 7% of the total number of samples assayed for gold and 1% of the total number of samples assayed for silver.

Table 12-1: Summary of Analytical Quality Control Data Produced

between December 2011 and July 2012

Total

Au (%)

Comment

Sample Count

54,515

Field Blanks

640

640

1.2

Coarse marble

Certified Standards

2,648

2,648

4.9

P3B

449

449

CDN (0.409 ± 0.042 g/t Au)

P4A

241

241

CDN (0.438 ± 0.032 g/t Au)

435

435

CDN (0.946 ± 0.102 g/t Au)

279

279

CDN (0.972 ± 0.108 g/t Au)

1P5D

472

472

CDN (1.47 ± 0.15 g/t Au)

1P5E

144

144

CDN (1.52 ± 0.11 g/t Au)

CDN (4.77 ± 0.40 g/t Au, 101.8 ± 72.5 g/t Ag)

573

475

1,048

CDN (4.96 ± 0.42 g/t Au, 7.0 ± 4.8 g/t Ag)

Field Duplicates

524

524

1.0

Pulp Duplicates

Total QC Samples

3,812

3,812

7.0

Check Assays

At ALS Chemex

352

352

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The performance of the control samples is acceptable. For field blanks, assuming a threshold limit of five (5) times the detection limit, only one (1) sample performed above this threshold. The field blank chart does not show any evidence of sample contamination. For certified standards, the majority of samples performed as expected within two (2) standard deviations for both gold and silver. The certified standard charts do not show any evidence of analytical bias.

Paired data for field duplicates show that gold grades can be reasonably reproduced by ALS Chemex Laboratories. Rank half absolute difference (“HARD”) plots suggest that 62.8% of samples have HARD below 10%. This is consistent with results expected from gold mineralization from this deposit style. Quantile-quantile and mean-to-half relative deviation plots show no apparent bias between original and duplicate samples.

Paired data for check assays show that original gold assays produced by ALS Chemex can be reasonably reproduced by Actlabs. HARD plots suggest 77.3% of samples have HARD below 10%. Quantile-quantile and mean-to-half relative deviation plots show no apparent bias between original and check samples.

In the opinion of SRK, the results of the analytical quality control data produced from December 2011 to July 2012 are sufficiently reliable to support resource estimation.

12.3.3

Verification of Electronic Data

SRK visited the Emo, Ontario exploration office in March 2010 and undertook a random check of approximately 100 original assay records from hard copy assay certificates against data in the electronic assay database. SRK examined data from various periods and drill programs and found no significant errors in the Rainy River data, but noted a few discrepancies with the Nuinsco data. The hardcopy Nuinsco assay data were subsequently recaptured by an independent contractor from original hardcopy assay certificates.

Rainy River made available to SRK a Gemcom database containing all electronic data accumulated on the Rainy River Gold Project. SRK conducted a series of routine verifications to

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ensure the reliability of the electronic data provided by Rainy River. SRK validated all tables using Gemcom validation tools that check for gaps, overlaps and out of sequence intervals.

The Gemcom collar, survey, lithology and assay tables did not contain obvious errors. On completion of the validation procedures, SRK concludes that the digital database for the Rainy River Gold Project is reliable for resource estimation.

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13.

MINERAL PROCESSING AND METALLURGICAL TESTING

13.1

Historical Metallurgy

Initial metallurgical testwork was carried out at SGS in Lakefield, Ontario from 2008 to 2011 and was the basis for the PEA Update Technical Report (October 2012).

13.1.1

Sample Selection

The metallurgical samples for this testwork were selected to give an accurate representation of the different mineralization zones based on information available at the time.

A master composite sample was created by combining individual samples from each zone in the proportion indicated in Table 13-1.

Table 13-1: Master Composite Proportions


Zone Composite	Zone Composite Proportions			Total Proportion
CAP		2.0	%		20.0	%
Z-433		12.0	%
HS		1.0	%
NZ		5.0	%
ODM-1		35.1	%		80.0	%
ODM-2		3.5	%
ODM-3		31.4	%
ODM-4		9.9	%
Master					100.0	%

13.1.2

Historical Testwork

The testwork included mineralogy, comminution, gravity separation, flotation, flotation concentrate leaching and whole rock leaching. The results from the testwork indicated that the material was moderately hard. Overall gold recovery for the flotation concentrate leach circuit was estimated at 88.5% with the flotation feed ground to a P₈₀ of 150 µm and the flotation concentrate re-ground to a P₈₀ of 15 µm. Recovery for the whole rock leach circuit was estimated at 91.0% when ground to a P₈₀ of 60 µm. No optimization of the grind size was performed at the time.

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The flotation concentrate leach option was selected mainly due to the separation of the sulphides into a low mass stream that could be deposited separately from the flotation tailings which would be low sulphide and cyanide free. While this flowsheet was selected for the December 2011 PEA, there was no significant economic or environmental benefit to either option that justified a final flowsheet selection. For this reason, additional testwork was performed in 2012 to give the analysis more precision and allow for a more conclusive flowsheet selection.

13.2

Composite and Sample Selection

13.2.1

Composite Selection

The results presented in Sections 13.2 to 13.14 were used as the basis for the Feasibility Study. The testwork was completed from October 2011 until November 2012.

Three (3) composites were used for determining the flowsheet and the parameters for the variability testwork:

•

ODM Master;

•

Initial Pit; and

•

Remaining-Life-of-Mine (“RLOM”).

The ODM master composite was developed as the ODM Zone is the most significant portion of the deposit and Initial Pit. The initial pit composites and RLOM composites were selected to develop a better understanding of the metallurgical responses for the early years of mining compared to the later years.

The tonnages by zone that were used to develop the composites are presented in Table 13-2. The values shown are those used for metallurgical testing and were determined in March 2012.

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Table 13-2: Weight Percentages by Zone for Initial Pit and RLOM Composites and Overall Pit


	Initial Pit		RLOM		Overall Pit
Zone	(%)		(%)		(%)t
ODM		86.4		60.4		68.0
Z-433		4.3		13.8		11.0
HS		0.4		5.5		4.0
NZ		4.4		5.2		5.0
CAP		4.6		15.1		12.0

13.2.2

Variability Sample Selection

In order to provide a high level of confidence, 161 comminution and 208 leaching variability samples were selected for variability testwork. All samples selected for the comminution and leaching variability testwork were selected using a geometallurgical approach to provide good definition of the deposit.

The sample locations are presented in Figure 13-1.

LOGO

Figure 13-1: Sample Locations for Comminution (Left) and Leaching (Right) Variability Testwork

The sample selection provides a good representation of the entire proposed pit.

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It should be noted that a number of samples are located outside of the proposed pit outline. This is due to the change in size of the engineered pit from the December 2011 PEA to the August 2012 updated PEA. The pit was reduced in size, however, the testwork campaign had already commenced at this stage. This is illustrated by a pit cross section shown in Figure 13-2, where the August 2012 pit outline is in purple and the December 2011 pit is in grey.

LOGO

Figure 13-2: Sample Locations for Comminution (Left) and Leaching (Right)

Variability Testwork (Cross-Section View)

Figure 13-2 shows that the pit outline was reduced from the December 2011 PEA pit from which the testwork program was designed. While a number of samples are located outside the new pit, the samples still provide good coverage and definition of the deposit and the proposed pit.

13.3

Mineralogy

Gold deportment studies have been performed on samples from each zone during the 2011 and 2012 testwork campaigns, including: 5 ODM, 2 Z-433, 1 CAP, 1 HS and 1 NZ sample:

•

The samples were composed mainly of non-opaque minerals, with minor amounts of pyrite present (ranging from 2.5% in one (1) of the Z-433 Composites to 9.5% in the CAP composite).

•

Gold associated with pyrite can be recovered and concentrated using froth flotation;

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•

Gold encapsulated in pyrite is difficult to recover via cyanidation without a fine or ultra-fine grind;

•

Gold associated with the non-opaque minerals will be not be recovered by flotation but can be recovered by leaching of flotation tailings or by whole rock leach if the particle is not locked.

•

The gold occurs mainly as native gold, electrum and kustelite. Small amounts of Petzite (Ag₃AuTe₂) were also noted;

•

The gold occurs as liberated, attached and locked particles in most of the samples at a grind size of 150 µm except for the CAP sample;

•

Liberated and attached gold can be readily extracted with whole rock leaching at the grind size of 150 µm;

•

Locked gold will require finer grinding to be recovered by cyanide leaching. Locked gold represented the majority of the gold, indicating that these composites would require grinding finer than 150 µm prior to whole rock cyanidation to achieve a good gold recovery;

•

The CAP composite had gold occurring as locked inclusions in pyrite and non-opaque minerals only. This is an indication that the gold in the CAP composite will be difficult to recover by either flotation (due to gold being locked with non-opaque minerals) or whole rock leach (due to gold being locked in pyrite).

•

The gold grain size was relatively fine in all samples, with coarse gold (>100 µm) noted only in two (2) of the composites (HS and one of the Z-433 samples).

•

Overall, the two (2) Z-433 composites had the coarsest gold particles;

•

All other samples had fine gold grains, which the majority of the gold distribution falling into the <10 µm category;

•

The liberated gold tended to have the coarsest grain size whereas the locked gold tended to be the finest.

•

Trace amounts of pyrrhotite were noted in approximately half the samples.

•

Pyrrhotite can have a negative impact on gold dissolution and can increase cyanide consumption.

13.4

Flowsheet Selection Testwork

Subsequent testwork was carried out from fall 2011 to spring 2012 to provide a better definition to the flowsheets and allow for a final flowsheet selection.

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Two (2) flowsheet options were investigated:

•

Flotation Concentrate Leach: This flowsheet includes rougher and cleaner flotation in which the sulphides are separated from gangue. The flotation concentrate is reground and the gold and silver are then recovered from the flotation concentrate using conventional cyanide leaching and carbon-in-pulp (“CIP”). A gravity circuit is included in the grinding circuit prior to flotation;

•

Gravity Tailings Leach: This flowsheet is a whole rock cyanide leach process followed by CIP. A gravity circuit is included in the grinding circuit prior to leaching.

Testwork was performed to give a better definition of both flowsheets. Once testwork was completed, a techno-economic study was performed to determine which process had the best economic return. All testwork was performed on an ODM master composite. This composite was selected since the ODM represents the largest portion of the deposit and Initial Pit.

13.4.1

Flotation Option

Flotation

The use of a flotation circuit prior to leaching was investigated. In this processing option, gravity gold is recovered in a gravity separation stage and the gravity tailings are subjected to froth flotation. The concentrate from the flotation circuit is reground and then leached while the tailings are discarded. The rougher flotation results are presented in Table 13-3.

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Table 13-3: Rougher Flotation Results


Test No.	Combined Products K₈₀, µm		Product	Mass %		Assays						% Distribution
						Au (g/t)		Ag (g/t)		S⁼ (%)		Au				Ag				S⁼
						Au (g/t)		Ag (g/t)		S⁼ (%)		Flot (unit)		Grav+ Flot		Flot (unit)		Grav+ Flot		S⁼
1		163	Ro Conc 1-5		14.6		4.33		18.3		12.9		91.9		94.3		77.6		78.7		97.4
			Rougher Tail		85.4		0.07		0.9		0.06		8.1		5.7		22.4		21.3		2.6
			Head (calc)		100.0		0.69		3.4		1.94		100.0		100.0		100.0		100.0		100.0
2		101	Ro Conc 1-5		14.6		4.03		19.5		13.0		92.0		94.3		78.7		79.7		97.8
			Rougher Tail		85.4		0.06		0.9		<0.05		8.0		5.7		21.3		20.3		2.2
			Head (calc)		100.0		0.64		3.6		1.94		100.0		100.0		100.0		100.0		100.0
3		79	Ro Conc 1-5		20.6		2.70		13.7		9.3		90.9		93.6		78.1		79.1		97.2
			Rougher Tail		79.4		0.07		1.0		0.07		9.1		6.4		21.9		20.9		2.8
			Head (calc)		100.0		0.61		3.6		1.98		100.0		100.0		100.0		100.0		100.0
4		76	Ro Conc 1-5		19.6		3.38		14.6		10.2		92.7		94.8		79.9		80.8		98.0
			Rougher Tail		80.4		0.07		0.9		<0.05		7.3		5.2		20.1		19.2		2.0
			Head (calc)		100.0		0.71		3.6		2.04		100.0		100.0		100.0		100.0		100.0

The rougher flotation results yielded gold recoveries from the gravity tailings ranging from 90.9 to 92.7% with mass pulls ranging from 14-20%. The sulphide recoveries were high, ranging between 97-98%. The high sulphide recovery is an indication that the gold flotation recovery likely cannot be improved significantly from these results since the gold is recovered with the sulphides. The silver recoveries in flotation ranged from 78.7 to 80.8%.

The use of a cleaner stage (to process the rougher concentrate) was investigated to reduce mass pull and increase gold grade in the feed to the cyanide leaching circuit. The results for the rougher and cleaner stages are presented in Table 13-4.

For the first test the concentrate was reground. In the second test, the concentrate was floated without regrinding.

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Table 13-4: Cleaner Flotation Results


Feed Size P₈₀, µm	Product	Mass %		Assays						% Distribution
				Au (g/t)		Ag (g/t)		S⁼ (%)		Au				Ag
				Au (g/t)		Ag (g/t)		S⁼ (%)		Flot		Grav+ Flot		Flot		Grav+ Flot		S⁼
30	1st Cleaner Conc		6.2		11.1		36.8		27.9		88.2		91.7		69.1		70.5		93.4
	1st Cleaner Conc + Scav		7.1		9.88		33.3		25.1		88.9		92.2		70.8		72.1		95.2
	Rougher Conc		17.2		4.15		15.0		10.4		91.0		93.6		77.5		78.6		96.4
	Rougher Tail		82.8		0.09		0.9		0.08		9.0		6.4		22.5		21.4		3.6
	Head (calc)		100		0.78		3.3		1.86		100		100		100		100		100
59	1st Cleaner Conc		6.7		8.91		36.3		0.8		81.7		87.1		66.7		68.2		2.1
	1st Cleaner Conc + Scav		7.4		8.58		34.5		1.2		86.7		90.6		69.9		71.2		3.4
	Rougher Conc		17.1		3.92		16.9		15.0		92.0		94.3		79.5		80.4		96.9
	Rougher Tail		82.9		0.07		0.9		0.10		8.0		5.7		20.5		19.6		3.1
	Head (calc)		100		0.73		3.6		2.65		100		100		100		100		100

The results indicated that the mass pull could be decreased to below 10% with a drop in gold recovery between 2% and 10%. Also, regrinding appeared to improve the recovery with a slightly lower mass pull. Losses in the silver recoveries were slightly more pronounced, ranging from 8 to 10%.

Flotation Concentrate Leach Tests

Cyanide leaching was performed on the flotation concentrate to estimate the recovery of the entire circuit. The product from the flotation tests was reground to two (2) particle sizes (approximately 50 and 15 µm) to investigate the effect of grind size on the cyanidation recovery. Bulk gravity and flotation separations were performed to generate the flotation concentrate for these tests.

The flotation concentrate leach results are presented in Table 13-5 and Table 13-6.

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Table 13-5: Flotation Concentrate Leaching Results (Gold Assays)


Number of tests	Test Parameters					Reagent Consumptions (kg/t)				Au Recovery (%)												Au Assays
Number of tests	P₈₀ (µm)		Pb(NO₃)₂ Addition (g/t)		O₂ or Air	NaCN		CaO		24 h		36 h		48 h		Grav		Float		Grav + Float + Leach		Residue (g/t)		Float Head Grade (g/t)		Calc. Head Grade (g/t)
3		45		0	Air		2.8		1.2		87.4		86.0		87.7		36.6		54.7		84.6		0.51		4.2		1.12
3		42		0	O₂		0.6		1.6		88.5		87.7		88.8		36.6		54.7		85.2		0.50		4.3		1.14
3		45		500	Air		2.7		1.3		88.1		89.0		86.7		36.6		54.7		84.0		0.55		4.2		1.12
3		46		500	O₂		0.6		1.4		87.7		86.9		87.5		36.6		54.7		84.5		0.51		4.1		1.10
3		14		0	Air		3.8		3.9		96.2		95.1		94.9		36.6		54.7		88.5		0.21		4.1		1.11
3		15		0	O₂		1.8		1.6		94.9		94.8		93.1		36.6		54.7		87.5		0.30		4.1		1.11
3		13		500	Air		4.4		3.5		95.8		94.8		94.1		36.6		54.7		88.1		0.21		3.9		1.06
3		14		500	O₂		2.3		1.9		95.9		94.9		94.4		36.6		54.7		88.3		0.21		4.1		1.11

Table 13-6: Flotation Concentrate Leaching Results (Silver Assays)


Number of tests	Test Parameters					Reagent Consumptions (kg/t)				Ag Recovery (%)												Ag Assays
Number of tests	P₈₀ (µm)		Pb(NO₃)₂ Addition (g/t)		O₂ or Air	NaCN		CaO		24 h		36 h		48 h		Grav		Float		Grav + Float + Leach		Residue (g/t)		Head Grade (g/t)		Calc. Head Grade (g/t)
3		45		0	Air		2.8		1.2		67.4		68.0		69.5		1.6		74.6		53.4		5.90		16.9		3.65
3		42		0	O₂		0.6		1.6		71.0		72.2		75.5		1.6		74.6		57.9		5.05		18.8		3.61
3		45		500	Air		2.7		1.3		67.2		70.1		68.7		1.6		74.6		52.8		6.10		19.1		3.65
3		46		500	O₂		0.6		1.4		71.1		69.8		70.7		1.6		74.6		54.3		5.40		18.0		3.67
3		14		0	Air		3.8		3.9		82.4		83.9		89.3		1.6		74.6		68.2		1.95		18.5		3.56
3		15		0	O₂		1.8		1.6		86.5		90.4		89.1		1.6		74.6		68.1		2.10		18.7		3.59
3		13		500	Air		4.4		3.5		83.4		88.5		89.8		1.6		74.6		68.6		2.00		18.8		3.60
3		14		500	O₂		2.3		1.9		89.7		90.5		94.0		1.6		74.6		71.7		1.20		19.0		3.64

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The results indicated that gold recoveries of 94 to 96% in the leaching circuit could be achieved by grinding the material down to around 15 µm. This gave overall recoveries ranging from 87.5 to 89.2% for the entire circuit. Silver recoveries were also higher when ground to around 15 µm, with recoveries ranging from 89.1 to 94.0%, compared to 68.7 to 75.5% when ground to approximately 45 µm.

High cyanide (NaCN) and lime (CaO) consumptions were noted when the samples were ground to 15 µm or below, ranging from 4.0 to 5.0 kg_NaCN/t and 2.0 to 6.0 kg_CaO/t. The cyanide and lime consumptions were considerably lower when ground to 40-50 µm, ranging from 1.8 to 4.9 kg_NaCN/t. The use of oxygen instead of air decreased the consumption of cyanide and lime to around 1.3 to 2.5 kg_NaCN/t and 1.3 to 2.0 kg_CaO/t for the samples ground to 15 µm. The use of lead nitrate was also investigated but did not have any noticeable effect.

Flotation Tailings Leaching

Additional testwork was performed in an attempt to improve the recovery of the circuit by leaching the flotation tailings.

The results from the flotation tailings leaching tests are presented in Table 13-7.

Table 13-7: Flotation Tailings Leaching Results


CN Test No.	Feed (Tails from Test)	Feed Size P₈₀, µm		Reag. Consumption kg/t of CN Feed				% Au Recovery										Residue Au (g/t)		Head Au (g/t)
				Reag. Consumption kg/t of CN Feed				CN Leach Time (h)								Rec.¹
				NaCN		CaO		6		24		36		48		Rec.¹
CN-8	F-1		181		0.08		0.35		66.6		66.8		67.0		67.3		3.9		0.02		0.06
CN-9	F-2		101		0.06		0.33		66.5		66.7		66.9		67.3		3.8		<0.02		0.06
CN-10	F-3		81		0.08		0.33		66.4		66.7		67.0		67.3		4.3		0.02		0.06
CN-11	F-4		82		0.08		0.43		66.5		66.7		66.9		67.2		3.5		0.02		0.06

¹	Recovery values indicates additional overall recovery.

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Based on the results, it can be seen that close to an additional 4% overall recovery could likely be achieved by leaching the flotation tailings. The recoveries presented in Table 13-7 are artificially high as they do not include the flotation concentrate leaching recoveries. Regardless, this option was rejected due to the high capital and operating costs associated with leaching of the flotation tailings.

Ultrafine Grinding

Ultrafine grinding tests were performed to simulate both an IsaMill and a stirred media detritor (“SMD”). The IsaMill tests were performed in a single unit to grind a flotation concentrate sample from an F₈₀ of 127 µm to a target P₈₀ of 10 µm. This target was selected prior to the completion of the flotation concentrate leach tests presented previously in Section 13.4.1. The tests were performed in a single IsaMill M4 test unit using 3.5 mm media. The results from the testwork are presented in Table 13-8.

Table 13-8: IsaMill Testwork Results


Raw Data
Solids SG: 3.06					Media Start (g): 6124						Media End (g): 6089.4
Calculated Data
Pass #	Gross kW		Net kW		Q (m³/h)		% Solids		M (t/h)		E (kWh/t)		Cumul. E (kWh/t)		P₉₈ (µm)		P₈₀ (µm)
Feed		—		—		0.158		44.6		0.101		—		0		459.8		127.5
1		1.76		1.00		0.138		44.7		0.088		11.3		11.3		97.6		26.0
2		1.69		0.92		0.138		45.8		0.092		10.0		21.3		52.0		22.3
3		1.69		0.93		0.138		46.9		0.095		9.7		31.0		41.8		19.5
4		1.68		0.92		0.138		45.9		0.092		10.0		41.0		37.9		18.4
5		1.67		0.90		0.136		45.9		0.092		10.0		51.0		35.4		17.5
6		1.69		0.93		0.14		45.9		0.093		10.0		61.0		33.9		16.8
7		1.68		0.92		0.138		45.9		0.091		10.0		71.0		32.9		16.3
8		1.70		0.93		0.091		45.5		0.059		15.7		86.7		30.9		15.4
9		1.60		0.83		0.095		40.9		0.054		15.5		102.2		29.7		14.8
10		1.60		0.83		0.098		41.1		0.056		15.0		117.2		28.9		14.3
Target Size (µm):				10		Est. kWh/t to Target:				472.4		Media Consumption (g/kWh):						21.6

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The results from the IsaMill tests indicated that between 80-100 kWh/t would be required to grind to 15 µm and the target grind size of 10 µm could not be achieved using the 3.5 mm media. An extrapolation of the results indicated that 472.4 kWh/t would be required to grind to 10 µm. Media consumption was estimated to be 21.6 g/kWh or 10.2 kg per tonne of flotation concentrate. The results are indicative of a material that is difficult to grind to 15 µm or below. The results from the test are plotted in Figure 13-3.

LOGO

Figure 13-3: Specific Energy vs. Particle Size for IsaMill Test

Testing at Metso was also performed to estimate the energy requirement of an SMD unit. The SMD tests were performed in a single unit to grind a flotation concentrate sample from an F₈₀ of 34 µm to a target P₈₀ of 10 µm. The tests were performed in a steel jar rotating at 76% critical speed using 19 mm steel balls. The particle sizes of the feed and products were analyzed using a Malvern Mastersizer 2000.

The results are presented in Table 13-9.

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Table 13-9: Stirred Media Detritor Testwork Results


	Pass # (% Passing)
Particle Size (µm)	Feed		1		2		3		4
589		100		100		100		100		100
417		99.2		99.7		100		100		100
295		97.6		99.5		100		100		100
208		95.5		99.4		100		100		100
147		93.3		99.4		100		100		100
104		90.4		98.4		99.4		100		100
74		87.4		97.0		98.5		100		100
53		84.9		96.0		98.2		100		100
44		83.4		95.4		98.2		100		100
37		81.6		94.4		98.1		100		100
25		74.4		88.6		95.0		99.6		100
18		64.2		78.4		87.0		95.8		98.9
12.5		49.6		62.0		71.5		84.7		91.3
8.8		35.2		44.6		53.2		67.7		77.1
6.3		23.5		30.1		36.6		49.6		59.5
4.4		15.0		19.5		23.9		33.7		42.0
3.1		9.4		12.5		15.4		22.1		28.0
P₉₈ (um)		325.5		95.4		36.6		22.1		17.3
P₈₀ (um)		34.1		19		15.4		11.4		9.5

Specific Energy (kWh/t)		0		20		40		80		120

It was estimated that approximately 45 kWh/t would be required to grind to 15 µm and 110 kWh/t to grind to 10 µm. The results confirmed that the flotation concentrate is difficult to grind to below 20 µm. A graphical representation of the energy requirement is presented in Figure 13-4.

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LOGO

Figure 13-4: Specific Energy vs. Particle Size for SMD Test

13.4.2

Gravity and Gravity Tailings Whole Rock Leach

A gravity circuit was included in the whole rock leach flowsheet. The gravity circuit is designed to remove and recover any gold nuggets which would not be completely leached in conventional cyanide leaching circuit. The use of a gravity circuit also allows for a more consistent head grade in the leach feed.

Leaching tests were performed on gravity tailings to compare a gravity tailings leaching circuit to the flotation concentrate leaching option. All tests were performed on an ODM master composite. The tests were done at grind sizes ranging from 50 to 120 µm.

The gravity tailings leaching results for gold and silver are presented in Table 13-10 and Table 13-11, respectively. Any results presented for grind sizes with more than one (1) test are averaged values.

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Table 13-10: Gravity Tailings Leach Results (Gold)


Number of tests	P80 (µm)		Reagent Consumptions (kg/t)				Au Recovery (%)												Au Assays
Number of tests	P80 (µm)		NaCN		CaO		6 h		24 h		36 h		48 h		Grav		Grav. + CN		Residue (g/t)		Head Grade (g/t)
1		119		0.08		0.39		77.7		83.5		85.6		85.8		29.1		89.9		0.10		0.98
1		95		0.12		0.38		76.9		86.3		85.3		86.7		29.1		90.6		0.10		0.98
3		68		0.16		0.39		78.7		88.3		84.6		89.3		29.1		92.4		0.08		0.98
1		50		0.34		0.40		79.2		87.9		88.4		89.8		29.1		92.8		0.08		0.98
3		94		0.09		0.34		77.2		87.6		87.0		88.1		25.7		91.1		0.10		1.05
2		75		0.10		0.31		79.3		89.9		87.8		90.1		25.7		92.6		0.08		1.05
4		62		0.14		0.36		79.3		87.5		88.1		89.6		25.7		92.3		0.08		1.05
3		51		0.18		0.37		78.8		90.6		88.6		90.8		25.7		93.2		0.07		1.05

Table 13-11: Gravity Tailings Leach Results (Silver)


Number of tests	P80 (µm)		Reagent Consumptions (kg/t)				Ag Recovery (%)												Ag Assays
Number of tests	P80 (µm)		NaCN		CaO		6 h		24 h		36 h		48 h		Grav		Grav. + CN		Residue (g/t)		Head Grade (g/t)
1		119		0.08		0.39		53.4		61.5		64.7		66.5		4.6		68.0		1.20		3.80
1		95		0.12		0.38		54.5		64.5		67.3		68.9		4.6		70.3		1.10		3.80
3		68		0.16		0.39		54.4		64.4		63.2		68.8		4.6		70.2		1.13		3.80
1		50		0.34		0.40		52.9		63.5		65.7		68.0		4.6		69.5		1.20		3.80
3		94		0.09		0.34		60.8		70.6		73.0		74.7		6.7		76.4		0.87		3.80
2		75		0.10		0.31		72.0		84.9		83.5		85.8		6.7		86.7		0.40		3.80
4		62		0.14		0.36		59.9		69.6		72.3		72.8		6.7		74.6		0.96		3.80
3		51		0.18		0.37		56.2		67.1		68.6		71.6		6.7		73.5		1.05		3.80

Test results showed gold recoveries ranging from 89.2 to 93.5% (residues from 0.07 to 0.12 g/t) and silver recoveries ranged from 67.2 to 81.0% (residues of 0.7 to 1.2 g/t). Sodium cyanide consumptions averaged 0.10 kg_NaCN/t, while lime consumptions averaged 0.36 kg_CaO/t. It was also noted that cyanide consumption increased with decreasing grind size.

From these results, it was estimated that an overall gold recovery of approximately 92% could be achieved when the material was ground below 90 µm followed by gravity tailings leaching.

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13.4.3

Flowsheet Selection

Based on a trade-off study comparison, it was determined that the whole rock leaching option with gravity separation was the most economically viable alternative and was therefore used as the basis for the Feasibility Study. The main reason for this selection was the significant amount of energy associated with regrinding the flotation concentrate and the high cyanide consumption in the flotation concentrate leaching, in addition to the risk associated with ultrafine grinding of this material. All subsequent testwork was therefore based on cyanide leaching of the gravity tailings.

13.5

Comminution Tests

A large comminution testwork program was undertaken to establish the basis for the sizing of the crusher, SAG mill and ball mill. The tests included 21 crushing work index tests (seven (7) tests at three (3) separate testing facilities), 16 Bond Ball Mill Work index (“BWi”) tests, 160 ModBond Ball Mill Work index (“ModBWi”) tests, 13 JK Drop Weight tests (“DWT”), 175 SAG Mill Comminution (“SMC”) tests and eleven (11) SPI tests. In addition, seven (7) samples were sent to Starkey and Associates for SAGDesign testing.

13.5.1

Crusher Work Index

Crushing work index (“CWi”) tests were performed at three (3) testing facilities (SGS and two (2) suppliers). The results from the testwork are presented in Table 13-12.

Table 13-12: Crusher Work Index Results


Lab	SGS/Phillips										Supplier A								Supplier B
Zone	ODM		Z-433		HS		NZ		CAP		ODM		Z-433		HS		CAP		ODM		Z-433		HS		CAP
No. of Tests		4		2		1		1		1		6		1		2		2		4		1		1		1
No. of Specimens		69		38		20		17		20		60		10		20		20		40		10		10		10
Average		19.7		34.8		25.0		19.4		10.9		20.9		18.7		18.8		14.0		11.6		10.3		10.3		7.3
Minimum		8.8		17.2		17.1		13.7		6.6		11.1		12.0		10.0		10.2		2.9		6.4		6.9		3.7
Median		17.7		35.9		24.5		17.4		10.1		20.9		18.2		17.3		14.3		10.3		10.1		9.8		6.7
80^thPercentile		24.0		39.9		28.4		24.4		14.3		24.7		23.4		21.7		15.5		16.5		11.1		13.4		9.8
Maximum		52.1		50.3		30.9		27.6		18.3		36.6		27.8		39.4		20.0		30.2		20.9		15.1		11.6

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The SGS results indicated that the crusher work indices ranged from 10.1 to 35.9 kWh/t and 14.3 to 39.9 kWh/t at the 50^th and 80^th percentiles, respectively. It was noted that the Z-433 samples had much higher values while the CAP Zone had lower CWi values. Supplier A produced CWi values ranging from 10.3 to 20.9 and 15.5 to 24.7 kWh/t at the 50^th and 80^th percentiles, respectively. The numbers provided by Supplier B were considerably lower, with values ranging from 6.7 to 14.3 kWh/t and 9.8 to 16.5 kWh/t at the 50^th and 80^th percentiles. The results from SGS and Supplier A are indicative of a hard material resistant to coarse breakage, while Supplier B indicated that the samples had average resistance to coarse breakage.

The results from Supplier A were used for the design as they were deemed the most reliable and consistent. An 80^th percentile value of 25 kWh/t was chosen for the design.

13.5.2

Unconfined Compressive Strength

Unconfined compressive strength tests were performed at Queen’s University to determine competency of the selected rock samples. Seven (7) samples (four (4) ODM, one (1) Z-433, one (1) HS and one (1) CAP) were tested with one (1) repeat each.

Ten (10) of the 14 samples had partial failure occur along foliation lines, including all of the ODM samples. The values from all the tests ranged from 34.5 to 109.4 MPa with an average of 66.3 MPa. The average of the samples that did not have partial failures along the foliation lines was 87.2 MPa. As most of the samples suffered partial failures along foliation lines, the results are not considered to be reliable and were not used for the design.

13.5.3

JK Drop Weight and SAG Mill Comminution

A large SAG Mill sizing testwork campaign including 13 JK DWT tests and 175 SMC tests was undertaken at SGS. The JK DWT tests were performed on PQ core drilled specifically for the comminution program, while the SMC tests were performed on samples selected from the exploration drilling program. The SMC samples were selected using a geometallurgical approach to give a good definition of the deposit. The samples were selected by SGS geometallurgy group by dividing the deposit into one million tonne blocks (“domains”) and selecting a sample in each domain.

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The JK DWT results consist of A and b factors (often referenced as A x b) which indicate the resistance to impact breakage and a t_a value, which indicates the resistance to abrasion. The JK Drop Weight test results were also used to calibrate the SMC results. The calibration was performed by zone (i.e., the HS Zone JK DWT was used to calibrate the HS SMC samples). The SMC tests generate A and b factors similar to the JK DWT, along with M_ia, M_ic, M_ih and density values. The M_ia value is the coarse grinding work index, M_ic is the crushing work index and M_ih is the HPGR work index. All SMC tests were performed on the -22.4 /+19.2 mm size fraction.

An SMC test was performed on the remaining portion of each sample that was first tested using JK DWT. The objective of this was to provide confidence in the results obtained from the SMC tests and to support the validity of using SMC testing for the variability testwork. The results from the JK DWT are presented alongside the corresponding SMC results for the same sample in Table 13-13.

Table 13-13: JK Drop Weight (DWT) and Corresponding SMC Results


	JK DWT										SMC
Zone	A		B		A x b		t_a		r (g/cm³)		A		b		A x b		Axb % Difference
HS		76.4		0.30		22.9		0.32		2.79		75.4		0.33		24.9		8.7
HS		66.4		0.37		24.6		0.31		2.81		58.0		0.56		27.6		12.2
ODM		66.2		0.37		24.5		0.45		2.77		68.9		0.35		24.1		–1.6
		50.8		0.61		31.0		0.46		2.82		55.0		0.60		33.0		6.4
		54.9		0.55		30.2		0.48		2.83		54.2		0.64		34.7		12.9
		53.2		0.59		31.4		0.47		2.83		54.9		0.57		31.3		–0.3
		55.2		0.67		37.0		0.57		2.80		56.4		0.70		39.5		6.7
		50.0		0.79		39.5		0.43		2.75		60.8		0.65		39.5		0.0
CAP		67.0		0.37		24.8		0.35		3.02		58.6		0.45		26.4		6.4
CAP		59.5		0.40		23.8		0.21		2.92		79.1		0.34		26.9		13.0
Z-433		60.6		0.41		24.8		0.44		2.81		69.5		0.35		24.3		–2.0
Z-433		60.1		0.42		25.2		0.28		2.82		70.5		0.36		25.4		0.8
NZ		35.0		0.81		28.4		0.46		2.73		64.7		0.45		29.1		2.5

It can be seen from Table 13-13 that the calibrated SMC test results are slightly higher than the JK DWT results for the same sample. As the SMC test could only be performed once on the excess material, the minor difference confirmed that the SMC tests could be used for the variability analysis rather than performing a full JK DWT as a better distribution could be developed with

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multiple SMC tests. As can be seen in Figure 13-5, the JK DWT tests (indicated as diamonds) fall onto the curve generated by the multiple SMC tests, as expected.

LOGO

Figure 13-5: SMC Data Distribution with JK DWT Calibration Points

The distributions for the Z-433, HS and CAP Zones are quite narrow and consistent throughout each respective zone, while there is a wider range of values for the ODM and NZ Zones.

The multiple SMC test results provide definition of the hardness profile of the deposit.

The SMC A x b and M_ia results are presented in Table 13-14. It should be noted that a lower A x b or a higher M_ia number is indicative of a harder material.

Table 13-14: SAG Mill Comminution SMC and M_iaValues


Description	A x b												M_ia (kWh/t)
Zone	ODM		Z-433		HS		NZ		CAP		Waste		ODM		Z-433		HS		NZ		CAP		Waste
Number of Tests		95		19		12		21		26		2		93		19		11		21		25		2
Average		32.7		23.7		22.0		28.3		23.2		21.6		23.7		30.0		31.5		27.0		30.3		31.9
Minimum		62.6		38.6		24.9		63.3		34.7		22.0		13.8		19.9		28.5		13.5		21.6		31.4
Median		32.2		22.7		22.1		26.0		22.3		21.6		23.2		30.4		31.1		27.4		30.6		31.9
80^thPercentile		26.5		20.7		20.8		21.8		20.3		21.3		27.0		32.5		33.0		31.4		33.2		32.1
Maximum		20.7		19.0		19.0		20.0		18.0		21.1		32.8		35.6		35.2		34.6		37.4		32.3

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From this data, it can be seen that all zones tested are considered to be very hard. The ODM Zone is slightly less resistant to coarse breakage while the other zones and waste rock samples were considerably harder. This can be noted from both the A x b and M_ia values. The hardest zone noted in the testwork was the CAP Zone, with an A x b of 20.3 and a M_ia of 33.2 kWh/t at the 80^th percentile.

A x b values of 25.7 and 24.0 at the 80th percentile were estimated from JK DWT and SMC tests for the Initial Pit and remaining-life-of-mine, respectively, using the proportions from Table 13-2. The A x b and ta values were used in the JKSimMet simulation program to estimate SAG mill sizing and energy requirements.

13.5.4

SAGDesign

Starkey and Associates (www.sagdesign.com) performed “SAGDesign” testwork on the PQ core samples that were tested using the JK DWT method. The testwork results are presented in Table 13-15.

Table 13-15: SAGDesign Testwork Results


Sample		SAG Mill Data from SAGDesign Test										Ball Mill Data from SAGDesign Test
No.	Name	Charge Mass (g)		No. of Revs		Bulk SG (g/cm³)		SG Solids (g/cm³)		Calc W_SAG to 1.7mm (kWh/t)		SAG Dis. Bond BWI (kWh/t)		Calc W_BM to P₈₀ (kWh/t)		Total W_T to P₈₀ (kWh/t)
1	NR 11-935		6906		1541		1.53		2.76		11.4		11.1		10.1		21.5
2	NR 11-955		7016		1292		1.56		2.80		9.5		11.0		10.0		19.5
3	NR 11-956		7145		1333		1.59		2.86		9.7		13.5		12.3		22.0
4	NR 11-1003		7205		1512		1.60		2.79		10.9		11.5		10.5		21.4
5	NR 12-1176		7499		1658		1.67		2.97		11.6		14.4		13.2		24.8
6	NR 11-825		6759		1770		1.50		2.74		13.3		14.8		13.5		26.8
7	NR 12-1191		6827		1626		1.52		2.77		12.2		14.6		13.3		25.4
Min			6759		1292		1.50		2.74		9.5		11.0		10.0		19.5
Median			7016		1541		1.56		2.79		11.4		13.5		12.3		22.0
Average			7051		1533		1.57		2.81		11.2		13.0		11.8		23.1
Max			7499		1770		1.67		2.97		13.3		14.8		13.5		26.8
Design Data (80^thPercentile)									2.81		12.2		14.6		13.3		25.4

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The W_SAG values provided by SAG Design ranged from 9.5 to 13.3 kWh/t with an 80^th percentile of 12.2 kWh/t.

Bond Ball Mill Work index tests were performed on the discharge of the SAGDesign testwork at a closed side setting of 200 mesh. The design BWI value was determined to be 14.6 kWh/t.

The samples tested using the SAGDesign method were also tested by JKDWT, presented in Section 13.5.3. This allowed for a direct sample-by-sample analyses of the SAGDesign method compared to the JKDWT method.

A comparison of the SAGDesign and JK DWT test program results is presented in Table 13-16.

Table 13-16: Comparison between SAGDesign and JK DWT Results


Sample Name	SAGDesign (kWh/t)		JK DWT (A x b)		JKSimMet (kWh/t)
NR 11-935		11.4		24.5		13.6
NR 11-955		9.5		31.0		11.9
NR 11-956		9.7		30.2		12.1
NR 11-1003		10.9		31.4		11.8
NR 12-1176		11.6		24.8		13.4
NR 11-825		13.3		22.9		13.9
NR 12-1191		12.2		25.2		13.3

The SAGDesign results confirmed the JK Drop Weight results and correlated well in terms of hardness. It can be seen that the softer samples according to the JK DWT (NR 11-955, NR 11-956 and NR 11-1003) had the lowest pinion power requirements according to SAGDesign (ranging from 9.5 to 10.9 kWh/t). Furthermore, the hardest sample from the JK DWT (NR 11-825) had the highest pinion power requirements (13.3 kWh/t). The overall results indicate that the SAGDesign method yielded pinion energy requirements that were approximately 1-2 kWh/t lower than the JKSimMet method. This good correlation between the two (2) methodologies provides confidence in the sizing of the SAG mill.

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13.5.5

Bond Work Index

The ball mill sizing test program consisted of 160 Modified Bond Mill Work index (“ModBWi”) tests and 20 standard Bond Ball Mill index (“BWi”) tests. The tests were performed on all five (5) zones of the deposit.

To validate the ModBond numbers, a comparison between the ModBond Ball Mill Work index and the full Bond Ball Mill Work index was performed. The average ModBond and Bond Ball Mill Work indices were identical for the two (2) tests, and overall the tests indicated that the ModBond results were representative of the results obtained by the full Bond Ball Mill tests.

A visual representation of the results was prepared to verify if the ModBWi were within 5% of the BWi. The results from the full BWi were plotted with a 5% error bar and the corresponding ModBWi results were also shown. The plot is presented in Figure 13-6.

LOGO

Figure 13-6: Full Bond Ball Mill Work Indices vs. ModBond Work Indices (200 Mesh)

It can be seen that all but four (4) of the ModBWi results fell within 5% of the BWi. Since the majority of the results were within 5% of the BWi and the average for the two (2) datasets was

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identical, the ModBWi were deemed representative of the BWi and the full program was performed using the ModBond technique.

The tests were performed at a closed side setting of 200 mesh (75 µm). The results are presented in Table 13-17.

Table 13-17: Bond and ModBond Results


	Bond Work Index (200 mesh)										ModBond Work Index (200 mesh)
Description	kWh/t										kWh/t
Zone	ODM		Z-433		HS		NZ		CAP		ODM		Z-433		HS		NZ		CAP
Number of Tests		5		4		2		2		3		89		17		10		20		24
Average		13.6		15.6		16.2		13.0		15.2		13.8		15.1		14.9		14.1		14.7
Minimum		12.6		15.2		16.1		12.1		14.8		11.6		12.9		14.1		11.1		13.0
Median		13.8		15.7		16.2		13.0		14.9		13.8		15.3		15.0		14.2		14.8
80^thPercentile		14.2		15.9		16.2		13.5		15.6		14.7		15.4		15.2		15.0		15.5
Maximum		15.0		15.9		16.3		13.8		16.1		16.0		15.8		15.5		16.2		15.8

It can be seen that at the 80^th percentile all the zones are relatively similar in terms of ModBWi. A weighted average ModBWi value of 15.0 kWh/t (80^th percentile) for the deposit was determined using the ModBond test results.

The distribution of the ModBond work indices can be seen in Figure 13-7.

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LOGO

Figure 13-7: Distribution of ModBond Indices (200 Mesh) by Zone

The ODM and NZ Zones were slightly softer than the other zones and also had a wider range of values, analogous to the SMC tests. However, at the 80^th percentile the BWi values are much closer across all the zones ranging from 14.7 kWh/t to 15.5 kWh/t. A weighted average value of 15.0 kWh/t was used for design.

13.5.6

ModBond and Axb

Figure 13-8 shows an interesting behaviour between A x b and ModBond. The linear trend indicates that the material grinding characteristics (impact and attrition) are very consistent (hard A x b with a hard BWi). This information may be useful in future production planning.

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LOGO

Figure 13-8: ModBond Wi vs. Axb

13.5.7

Bond Abrasion Index

Twenty-four abrasion index tests were performed and the results indicated a large amount of variability in the samples with values ranging from 0.050 to 0.663 (7^th to 88^th percentile hardness in the SGS database).

The results are presented in Table 13-18.

Table 13-18: Abrasion Index Results


Description	ODM		Z-433		HS		NZ		CAP
Number of Tests		12		4		2		2		4
Average (g)		0.20		0.27		0.32		0.11		0.15
Minimum (g)		0.05		0.14		0.21		0.11		0.08
Median (g)		0.15		0.21		0.32		0.11		0.15
80^thPercentile (g)		0.26		0.33		0.38		0.11		0.19
Maximum (g)		0.66		0.51		0.43		0.11		0.21

Overall, the ore is considered moderately abrasive.

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13.5.8

Comparison of Grinding Results from 2011 to 2012

The results from the 2012 testwork campaign were compared to those that were used as the basis of the December 2011 PEA. Considerably more SMC work was performed in 2012, allowing for a better definition of the hardness of the deposit.

A comparison of results obtained in 2011 and 2012 is presented in Table 13-19.

Table 13-19: SMC (Axb) Comparison of 2011 vs. 2012


	2011 Results						2012 Results
Description	ODM		Z-433		Other		ODM		Z-433		Other
Number of Tests		5		1		—		93		19		59
Average		38.2		31.0		—		32.7		23.7		24.7
Minimum		39.8		31.0		—		62.6		38.6		63.3
Median		39.2		31.0		—		32.2		22.7		23.0
80th Percentile		36.1		31.0		—		26.5		20.7		20.7
Maximum		35.8		31.0		—		20.7		19.0		18.0

It can be seen that far more samples were tested in the new campaign (171 vs. 6). Also, the results revealed that the deposit is considerably harder than was previously indicated from the 2011 results. In 2011, the ODM SMC values ranged from 35.8 to 39.8, with an 80^th percentile of 36.1. In 2012, the range was from 20.7 to 62.6 with an 80^th percentile of 26.5. The 2011 ODM samples corresponded to the 14^th to 28^th percentile of ODM results obtained in 2012.

The Z-433 sample tested in 2011 was also much softer than average for the Z-433 Zone, corresponding to the 5^th percentile of hardness from the 2012. It can also be seen that the other Zones (HS, NZ and CAP) are harder than the ODM Zone and that no testwork was available from these Zones in 2011.

Based on these results, the design Axb used for sizing the SAG mill was changed from 34.0 to 24.2.

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13.5.9

Grinding Circuit Design

Several different design methods were used to size the SAG and ball mill circuit. The 80^th percentile of the crushing and grinding testwork results was used for each of the tests to ensure that an appropriate safety factor was provided. The following estimation methods were used:

•

Morrell’s Equation: This method uses the M_iaand M_ic values from the SMC tests described in Section 13.5.3 and the Bond Ball Work index test results from Section 13.5.5 to estimate the total power required by the SABC circuit. While the power divided into the SAG mill, ball mill and pebble crusher, the total circuit energy is the recommended value to use.

•

JKSimMet with the Bond Equation: This method utilizes the A x b and T_aresults from the JK DWT and SMC tests to estimate the power requirement for the SAG mill using the software JKSimMet. The JKSimMet simulation also indicates the transfer size (T₈₀) which can then be used to size the ball mill according to the Bond Equation using standard efficiency factors and the BWI or ModBWi.

•

JKSimMet with the Phantom Cyclone: This method estimates the SAG mill power using the same method as previously described in the JKSimMet with the Bond Equation method. The SAG mill discharge is assuming that the SAG discharge is treated by a cyclone prior to feeding the ball mill circuit, with only the cyclone underflow being treated by the ball mill. The ball mill power is then calculated using the Bond Equation with no efficiency factors and the BWi or ModBWi. This method often yields a lower energy requirement than the standard Bond method with efficiency factors as it underestimates the hardness of the material in the cyclone underflow. It is BBA’s opinion that this method should not be used for design; however, it has been included in order to provide an additional reference.

•

SAGDesign: The testwork and interpretation performed by SAGDesign is described in Section 13.5.4.

•

Oreway Mineral Consultants (OMC): A third party was asked to independently assess the power requirements for the grinding circuit using the data presented in Section 13.5 excluding the SAGDesign testwork.

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To calculate the power requirement of the SAG and ball mill (at the pinion), the following design criteria were used:

•

Simulations performed at a nominal tonnage of 906 t/h or 20,000 tpd:

•

Energy requirements (operating work indices) were then used to determine the operating power and required installed power for the SAG mill and ball mill for a nominal tonnage of 21,000 tpd.

•

Variable transfer size (T₈₀), depending on method;

•

Final circuit P₈₀of 75 µm;

•

Axb value of 24.2, t_a value of 0.35;

•

BWi value of 15.0 kWh/t; and

•

M_iavalue of 29.3 kWh/t.

The results from the simulations are presented in Table 13-20.

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Table 13-20: SAG and Ball Mill Simulation Results


Method	Units	Morrell’s Equation		JKSimMet + Bond’s Equation		JK SimMet + Phantom Cy		SAGDesign		OMC
Method	Units	80^thPercentile		80^thPercentile		80^thPercentile		79^thPercentile¹		80^thPercentile
Parameters
F₈₀	µm		162 500		162 500		162 500		152 000		<150 000
T₈₀	µm		750		2 400		2 400		1 300		Unknown
Final P₈₀	µm		75		75		75		75		75
Energy Requirements (Operating Work Indices)
SAG Mill	kWh/t		15.26		13.23		13.23		12.56		13.70
Ball Mill	kWh/t		12.92		13.03		12.20		12.89		12.60
*Subtotal*	kWh/t		28.18		26.26		25.43		25.45		26.30
Pebble Crusher	kWh/t		0.46		0.37		0.37		—		0.57
*Total*	kWh/t		28.64		26.63		25.79		25.45		26.87
Operating Power Required (21,000 tpd)
SAG Mill	kW		14 510		12 580		12 580		11 948		13 033
Ball Mill	kW		12 289		12 395		11 603		12 262		12 143

¹	The 79^thpercentile used for the SAGDesign simulations was based on seven (7) samples only.

²	Simulations were performed at 20,000 tpd. Operating powers for 21,000 tpd were calculated using the same operating work index (kWh/t) as was used for the 20,000 tpd simulations.

Note: Particle sizes have not been adjusted for 21,000 tpd.

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It can be seen that the JKSimMet simulation using the Phantom Cyclone method yielded the lowest power requirement for the Ball mill, as expected. The lowest overall circuit (SAG mill + ball mill + crusher) energy requirement estimated was by SAGDesign at 25.45 kWh/t while Morrell’s Equation yielded the highest circuit energy requirement at 28.64 kWh/t. It can be seen that all the power requirements are relatively close, giving confidence to the calculation methods. The Feasibility Study design was based on the JK SimMet with the Bond Equation method which estimated a circuit energy requirement of 26.63 kWh/t. This matches well with the energy requirement calculated by OMC.

Based on these results, it is recommended that both the SAG mill and ball mill be equipped with 15 MW dual-pinion drives for a SAG feed throughput of 951 t/h. The SAG mill and drive were sized with design factors to provide sufficient operating flexibility to achieve the plant throughput. The ball mill drive size was selected to match the SAG drive for a reduction in spares and to simplify operations. BBA recommends that a 36’ x 20’ (18.25’ EGL) SAG mill and a 26’ x 40.5’ (40’ EGL) bBall mill be installed to draw this power, based on the equipment sizing software and discussions with mill suppliers. This design will allow for operational flexibility and provide confidence that the process plant throughput will be achieved.

13.6

Gravity Separation

13.6.1

Gravity Recoverable Gold

Two (2) Gravity Recoverable Gold (“GRG”) tests were performed on zone composites representing the ODM and Z-433 Zones. The test results are presented in Table 13-21.

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Table 13-21: Gravity Recoverable Gold Results


	Grind Size P₈₀ (µm)		Mass				Assay Au (g/t)		Au grams (g)		Dist’n %
Sample	Grind Size P₈₀ (µm)	Product	grams		%		Assay Au (g/t)		Au grams (g)		Dist’n %
ODM Master	650	Stage 1 Conc		79.1		0.4		46.7		3,691		18.8
	542	Sampled Tails		188.8		1.0		0.62		118		0.6
	275	Stage 2 Conc		76.0		0.4		48.7		2,698		18.8
	211	Sampled Tails		205.7		1.1		0.72		149		0.8
	141	Stage 3 Conc		98.6		0.5		27.1		2,676		13.6
	90	Final Tails		18,339		96.6		0.51		9.311		47.7
	Total (Head)			18,987		100		1.03		19,643		100
	Knelson Conc			254		1.3		39.7		10,065		51.2
Z433	612	Stage 1 Conc		80.2		0.40		56		4,529		20.6
	546	Sampled Tails		204.5		1.02		0.89		182		0.83
	260	Stage 2 Conc		87.9		0.44		52		4,568		20.8
	247	Sampled Tails		194.5		0.97		0.75		145		0.66
	132	Stage 3 Conc		109.8		0.55		35.8		3,927		17.9
	92	Final Tails		19,323		96.6		0.45		8,609		39.2
	Total (Head)			20,000		100		1.10		21,960		100
	Knelson Conc			277.9		1.39		46.9		13,024		59.3

The GRG numbers from these tests were 51.2 and 59.3, respectively. A higher GRG number indicates that more gold can be recovered by gravity; however, this cannot be taken as an absolute number. The actual gold recovery by gravity will be dependent on grind size and material flow to the gravity circuit.

13.6.2

Variability Gravity Testwork

In addition to the GRG tests, gravity separation was also performed in the variability test program using 2 kg samples. The gravity recoveries of these tests ranged from 1% to 77% with an average of 27% for the non-CAP Zones. The gravity gold recovery from the CAP Zone was considerably lower with an average recovery of 9%.

The gold and silver recoveries as a function of head grade are presented in Figure 13-9 and Figure 13-10.

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LOGO

Figure 13-9: Gold Gravity Recovery vs. Head Grade

LOGO

Figure 13-10: Silver Gravity Recovery vs. Head Grade

It can be seen that the gravity recovery of gold is slightly sensitive to the head grade. The same trend was noted for both the non-CAP and CAP Zone, however, the CAP gold recovery was lower than the non-CAP Zones. No trend was noted for the silver and it was assumed that silver gravity recovery is independent from the head grade. The silver gravity recovery for the CAP

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Zone was lower than the non-CAP Zones, analogous to the gold gravity recovery. The average silver gravity recovery for the CAP Zone was roughly 3% while the non-CAP Zones silver gravity recovery was approximately 10%.

13.7

Heap Leaching

Heap leaching was investigated as an alternative process to agitated tank leaching. The following conditions were used for the heap leaching tests:

•

Feed material: 12.7 mm (1/2 inch) material from ODM and Z-433 master composites;

•

50% pulp density;

•

0.5 g/L NaCN concentration; and

•

pH range: 10.5 – 11.

A visual representation of the results is presented in Figure 13-11.

LOGO

Figure 13-11: Heap Leach Gold Recovery Curve

The test results showed a 29.7% recovery after fourteen days with a plateau beginning around day five (5). Based on the low recovery, heap leaching was rejected as a processing option.

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13.8

Cyanide Leaching

13.8.1

Gravity Tailings Leaching

Gravity tailings leaching tests were performed on the Initial Pit and the remaining-life-of-mine (“RLOM”) composites. The tests were performed as “stop tests” with residue assays at each time duration.

Thirty-six tests were performed for each composite to help determine leach time and final grind size using the following criteria:

•

Four (4) leach times (18, 24, 30 and 36 h for the Initial Pit and 12 h, 18 h, 24 h and 30 h for the RLOM sample);

•

Three (3) grind sizes (110, 85 and 70 µm); and

•

Two (2) repeats (three (3) total tests) per leach time/grind size criteria.

The additional gravity tailings leaching test results are shown in Table 13-22 and Table 13-23 for gold and silver assays, respectively. The results are presented as averages of the 12 tests performed per grind size per composite.

Table 13-22: Additional Gravity Tailings Leaching Results (Gold Assays)


					Reagent Consumptions (kg/t)				Au Recovery (%)														Au Assays
Comp Name	Number of Tests		P₈₀ (µm)		NaCN		CaO		12 h		18 h		24 h		30 h		36 h		Grav		Grav + CN		Residue (g/t)		Head Grade (g/t)
Initial Pit		12		110		0.03		0.32		—		82.6		83.8		82.6		83.9		33.1		89.2		0.12		1.07
		12		85		0.04		0.33		—		84.8		86.1		85.2		86.4		33.1		90.9		0.10		1.07
		12		70		0.05		0.35		—		85.8		86.6		86.7		86.5		33.1		90.9		0.10		1.07
RLOM		12		110		0.02		0.31		79.7		79.7		81.3		82.2		—		29.6		87.5		0.10		0.83
		12		85		0.02		0.32		82.1		82.6		83.8		84.2		—		29.6		88.9		0.09		0.83
		12		70		0.01		0.32		84.1		85.2		86.0		85.7		—		29.6		90.0		0.08		0.83

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Table 13-23: Additional Gravity Tailings Leaching Results (Silver Assays)


					Reagent Consumptions (kg/t)				Ag Recovery (%)														Ag Assays
Comp Name	Number of Tests		P80 (µm)		NaCN		CaO		12 h		18 h		24 h		30 h		36 h		Grav		Grav + CN		Residue (g/t)		Head Grade (g/t)
Initial Pit		12		110		0.03		0.32		—		62.3		61.9		69.9		61.2		7.4		64.1		1.07		2.80
		12		85		0.04		0.33		—		59.4		62.4		70.5		62.8		7.4		65.5		1.07		2.80
		12		70		0.05		0.35		—		58.0		65.3		72.1		61.8		7.4		64.6		1.10		2.80
RLOM		12		110		0.02		0.31		61.1		66.4		68.1		68.9		—		6.1		70.8		0.80		2.80
		12		85		0.02		0.32		65.1		68.9		70.8		71.3		—		6.1		73.1		0.77		2.80
		12		70		0.01		0.32		66.2		72.1		70.7		70.8		—		6.1		72.6		0.80		2.80

The test results showed that the recovery of gold is moderately sensitive to grind size, confirming trends that had previously been noted. No noticeable improvement in recovery was noted beyond 24 hours of leaching for the Initial Pit; however, slight increases in recovery were noted for the RLOM sample when leaching from 24 to 30 h.

The gravity tailings residue for all the tests presented in Sections 13.4.2 and 13.8.1 were plotted versus the grind size of the feed material, presented in Figure 13-12. The dotted lines represent the sensitivity of the assay technique (0.02 g/t).

LOGO

Figure 13-12: Gravity Tailings Leach Residue vs. Grind Size

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To determine a final P₈₀ for the variability test program, a cost versus revenue study was performed. The costs included cyanide consumption, grinding energy at a fixed tonnage and estimated media wear, while the revenue was calculated based on the residue equation from Figure 13-12. High and low cost scenarios were investigated in addition to the nominal costs. The cost of sodium cyanide, steel and energy were varied to generate the high and low cost scenarios.

To standardize the values, the marginal costs and marginal revenues were used, rather than absolute costs and revenues, to determine the P₈₀ at which is it no longer economical to grind finer.

The results are presented in Figure 13-13.

LOGO

Figure 13-13: Cost and Revenue Analysis by Grind Size

The results show that for the average costs of the aforementioned parameters, grinding to 65 µm is still economical. However when using higher costs, it is only economical to grind to 75 µm.

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Based on these results, a grind size of 75 µm and a retention time of 36 h (with subsampling at 30 h) were selected for the variability test program.

13.8.2

Cyanide Concentration

The effect of the concentration of sodium cyanide in the leach was investigated by varying the sodium cyanide concentration from 0.15 to 0.5 g/L. The testwork was based on 36-hour cyanide leaching tests with subsampling at different time intervals.

The results from the cyanide concentration testwork results are presented in Table 13-24.

Table 13-24: Cyanide Concentration Testwork Results


					Reagent Consumptions (kg/t)				Au Recovery (%)														Au Assays
Comp Name	P₈₀ (µm)		NaCN Conc. (g/L)		NaCN		CaO		12 h		18 h		24 h		30 h		36 h		Grav		Grav + CN		Residue (g/t)		Head Grade (g/t)
RLOM		118		0.50		0.11		0.40		77		80		83		81		82.8		16.4		85.6		0.12		0.67
		117		0.30		0.08		0.37		71		77		82		81		81.9		16.4		84.9		0.13		0.69
		120		0.20		0.06		0.40		74		78		82		82		82.3		16.4		85.2		0.12		0.65
		118		0.15		0.06		0.41		70		77		81		80		82.3		16.4		85.2		0.12		0.68

The test results did not indicate that the gold or silver recoveries were sensitive to cyanide concentration.

The results are plotted in Figure 13-14.

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LOGO

Figure 13-14: Gold Recovery vs. Time at Different NaCN Concentrations

Based on these results, 0.5 g/L NaCN was selected for the variability tests as it did not appear that the cyanide level had any effect on overall recovery. This value was selected to ensure there was no cyanide starvation during variability testwork.

13.8.3

Pre-Conditioning

Pre-conditioning tests were performed on the sample to determine if aerating the sample prior to leaching had any effect on cyanide consumption. A short pre-conditioning was performed on the sample to raise the dissolved oxygen levels to approximately 5 ppm for four (4) tests while no pre-conditioning was performed for another four (4) tests. A level of 5 ppm dissolved oxygen is comparable to that of typical slurry after passing through the grinding circuit.

The pre-aeration results are presented in Table 13-25.

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Table 13-25: Pre-Conditioning vs. No Pre-Conditioning Testwork Results


				Reagent Consumptions (kg/t)				Au Recovery (%)												Au Assays
Comp Name	Pre- Aeration	P₈₀ (µm)		NaCN		CaO		6 h		12 h		24 h		36 h		Grav		Grav + CN		Residue (g/t)		Head Grade (g/t)
Initial Pit	Y		100		0.07		0.36		79		83		83		83.5		31.4		88.7		0.12		0.73
	Y		100		0.08		0.36		73		79		80		82.7		31.4		88.1		0.13		0.72
	N		100		0.22		0.30		74		82		86		84.0		31.4		89.0		0.12		0.72
	N		100		0.19		0.31		75		82		81		85.0		31.4		89.7		0.12		0.77
RLOM	Y		118		0.08		0.36		75		75		81		80.8		16.4		84.0		0.14		0.70
	Y		118		0.07		0.36		76		82		83		82.1		16.4		85.0		0.13		0.70
	N		118		0.18		0.33		72		77		80		81.5		16.4		84.5		0.13		0.70
	N		118		0.25		0.29		70		70		77		78.8		16.4		82.3		0.15		0.71

The cyanide consumption was increased from 0.07-0.08 kg/t in tests with pre-conditioning to 0.18-0.25 kg/t when no pre-aeration was performed. The residue levels when using pre-aeration were comparable to when pre-aeration was not used. Given the high consumptions when no-pre-conditioning was used, it was determined that cyanide would not be added to the grinding circuit. The use of a pre-conditioning tank is not required, as the minimum acceptable level of dissolved oxygen can be achieved through the grinding circuit.

Based on these results, it was decided that the variability tests would be performed with pre-conditioning prior to leaching.

13.8.4

Oxygen (O₂) vs. Air

The use of oxygen in the cyanide leach was investigated. The results from the tests did not indicate that the use of oxygen decreased cyanide consumption or improved reaction kinetics significantly.

The results are presented in Table 13-26.

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Table 13-26: O₂vs. Air and Lead Nitrate Addition Testwork Results


						Reagent Consumptions (kg/t)				Au Recovery (%)												Au Assays
Comp Name	Aera tion		Lead Nitrate	P₈₀ (µm)		NaCN		CaO		6 h		12 h		24 h		36 h		Grav		Grav + CN		Residue (g/t)		Head Grade (g/t)
		O₂	N				0.04		0.37		82		—		—		—		29.0		87.1		0.12
		O₂	N		54		0.04		0.36		—		86		—		—		29.0		90.2		0.09
		O₂	N		52		0.11		0.41		—		—		89		—		29.0		92.2		0.07
		O₂	N		61		0.06		0.38		—		—		88		—		29.0		91.8		0.10
		O₂	N		55		0.12		0.38		—		—		—		87		29.0		91.0		0.09
		O₂	N		59		0.04		0.39		—		—		—		87		29.0		90.8		0.10
Initial Pit		O₂	Y		59		0.16		0.50		—		—		—		88		29.0		91.5		0.08		0.97
		O₂	Y		45		0.05		0.52		—		—		—		87		29.0		90.9		0.08
		Air	Y		48		0.14		0.56		—		—		—		88		29.0		91.3		0.08
		Air	Y		59		0.06		0.51		—		—		—		87		29.0		90.9		0.09
		O₂	N		66		0.05		0.36		84		—		—		—		38.5		90.5		0.08
		O₂	N		59		0.05		0.41		—		87		—		—		38.5		91.9		0.07
		O₂	N		79		0.06		0.33		—		—		87		—		38.5		92.2		0.09
RLOM		O₂	N		68		0.07		0.40		—		—		84		—		38.5		90.3		0.08		0.89
		O₂	N		57		0.08		0.41		—		—		—		85.3		38.5		91.0		0.08
		O₂	N		66		0.08		0.41		—		—		—		85.5		38.5		91.1		0.08
		O₂	Y		70		0.06		0.53		—		—		—		84.0		38.5		90.2		0.08
		O₂	Y		71		0.03		0.53		—		—		—		84.6		38.5		90.5		0.08
		Air	Y		72		0.06		0.50		—		—		—		82.2		38.5		89.0		0.11
		Air	Y		71		0.08		0.49		—		—		—		84.1		38.5		90.2		0.10

Based on these results, it was decided that the variability tests would be performed with air.

13.8.5

Lead Nitrate Addition

The use of lead nitrate in the leaching circuit was also investigated. The results are presented in Table 13-26 in Section 13.8.4.

The lead nitrate cyanide leach test results did not indicate that the use of lead nitrate decreased cyanide consumption or improved reaction kinetics significantly. Furthermore, the lime addition for the samples with lead nitrate addition was higher than those without.

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Based on these results, it was decided that the variability tests would be done without lead nitrate.

13.8.6

Grade Recovery Variability Tests

Cyanide leaching was performed on 208 samples (and 37 repeats) and the results were used to develop grade-recovery curves for both gold and silver. The grade-recovery curves were based on the samples that met a set of criteria. All tests were performed at the following conditions:

•

Leach time of 36 hours with a subsample at 30 hours;

•

Target grind size (P₈₀) of 75 µm;

•

Cyanide concentration of 0.5 g/L NaCN;

•

30-minute pre-conditioning; and

•

pH of 10.5-11.

The summary of the variability testwork is presented in Table 13-27 and Table 13-28.

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Table 13-27: Gold Leaching Variability Testwork Average Results


							Reag. Consumption				% Au Recovery										Residue		Direct Au		Direct Au
	# of		P₈₀,		P₈₀ s³		(kg/t)				CN (Unit)								Overall		Au		Head		Head s³
Zone	Tests		µm		(µm)		NaCN		CaO		30 h		36 h		Grav		Overall¹		Recalculated²		(g/t)		(g/t)		(g/t)
ODM		138		95		40		0.06		0.37		78.1		78.7		25.8		83.8		90.1		0.12		1.19		1.84
Z-433		30		82		32		0.10		0.41		82.8		84.4		35.6		89.5		93.8		0.08		1.21		1.21
HS		13		86		26		0.06		0.36		84.4		86.1		24.2		89.1		90.8		0.05		0.51		0.31
NZ		24		86		33		0.08		0.40		82.1		82.7		27.0		87.0		91.2		0.07		0.75		0.81
*Non-CAP Subtotal*		*205*		91		38		*0.07*		*0.38*		*79.6*		*80.5*		*27.3*		*85.3*		*90.8*		*0.10*		*1.10*		*1.62*
CAP		40		92		50		0.11		0.62		71.5		71.5		8.7		73.9		76.8		0.16		0.67		0.74
TOTAL		245		91		40		0.08		0.42		78.3		79.0		24.3		83.5		89.3		0.11		1.03		1.52

¹	The overall recovery numbers were artificially low due to the large number of testwork samples and average residue values with low head grades.

²	Gold recovery recalculated using average direct head grade and average residue values.

³	One standard deviation s, indicates that 67% of the data falls within that deviation from the average value.

Table 13-28: Silver Leaching Variability Testwork Average Results


							Reag. Consumption				% Ag Recovery										Residue		Direct Ag		Direct Au
	# of		P₈₀,		P₈₀ s³		(kg/t)				CN (Unit)								Overall		Ag		Head		Head s³
Zone	Tests		µm		(µm)		NaCN		CaO		30 h		36 h		Grav		Overall¹		Recalculated²		(g/t)		(g/t)		(g/t)
ODM		138		95		40		0.06		0.37		57.4		58.8		10.0		62.7		67.1		1.24		3.77		7.38
Z-433		30		82		32		0.10		0.41		49.1		51.4		12.8		57.6		55.2		0.60		1.34		0.83
HS		13		86		26		0.06		0.36		47.8		48.2		8.5		52.8		51.1		0.51		1.04		0.69
NZ		24		86		33		0.08		0.40		55.8		56.1		8.5		59.5		62.2		0.53		1.39		0.72
*Non-CAP Subtotal*		*204*		91		38		*0.07*		*0.38*		*55.4*		*56.7*		*10.1*		*61.0*		*65.7*		*1.02*		*2.96*		*6.18*
CAP		40		92		50		0.11		0.62		63.8		65.1		3.0		66.4		67.5		0.86		2.65		1.25
TOTAL		244		91		40		0.08		0.42		56.8		58.1		9.0		61.9		66.0		0.99		2.91		5.68

¹	It should be noted that the overall recovery numbers are artificially low due to number of tests with low head grades or with residue grades below detection limit (0.5 g/t).

²	Silver recovery recalculated using average direct head grade.

³	One standard deviation indicates that 67% of the data falls within that deviation from the average value.

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The results indicated that there was significant variability in both the overall deposit and the individual zones. The average residue value for all the samples was 0.10 g/t Au for the zones other than CAP and 0.16 g/t for the CAP Zone. It can be seen that the CAP Zone had lower gravity and overall recoveries for gold, and lower gravity recoveries for silver.

It should be noted that the average overall recoveries for both gold and silver are lower than the true average that would be expected in operation. This is because the average overall recovery is more influenced by the samples with low head grades than the samples with average or high head grades. As a result, the overall recovery was recalculated using the average head grade and residue values, which is a more representative average recovery.

It can also be seen that the average P₈₀ of 91 µm was considerably coarser than the target P₈₀ of 75 µm. The P₈₀ also ranged from 30 to 260 µm, indicating that there were large variations in hardness of the material.

Filters were applied to the data to remove samples that did not conform to the process design criteria to generate the residue versus grade curves. Samples outside the proposed pit were removed, along with samples with a P₈₀ less than 60 µm and greater than 120 µm and those with high head grades (above 4 g/t Au or 20 g/t Ag). Samples with silver residues below the detection level of 0.5 g/t (unless the recovery was above 70%) and samples with gold recoveries below 60% were also removed.

The filtered gold and silver residues are plotted as a function of head grade in Figure 13-15 and Figure 13-16, respectively.

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LOGO

Figure 13-15: Gold Residue vs. Head Grade (Variability Tests)

LOGO

Figure 13-16: Silver Residue vs. Head Grade (Variability Tests)

Figure 13-15 shows that the CAP Zone has considerably higher gold residues than noted in the non-CAP Zones at the same head grade. No noticeable difference in silver residues was noted between the CAP and non-CAP Zones.

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13.8.7

Diagnostic Leach Testwork

Due to the lower gold recovery observed in the CAP Zone samples, and a small percentage of the non-CAP Zones samples, diagnostic leaches were performed on tailings from three (3) ODM and three (3) CAP samples. The objective of the diagnostic leach is to determine the nature of the residual gold from leach tests.

The diagnostic leach is composed of three (3) leaches:

•

Intensive Cyanide Leach: Extraction of gold that is readily available and is an indication that more retention time was required to complete reaction;

•

Hydrochloric Acid (HCl) Leach followed by Intensive Cyanide Leach: Extraction of gold that is associated with pyrrhotite, calcite, ferrites, etc. This is done by leaching the tailings using hydrochloric acid to dissolve the pyrrhotite and other minerals, then performing the intensive cyanide leach to extract the liberated gold; and

•

Aqua Regia Leach: Extraction of gold associated with or encapsulated by sulphide minerals such as pyrite and arsenopyrite.

The final residue from these leaches are considered to be locked in silicates or associated with fine sulphides that are locked in silicates.

The results from the leaches indicated that most of the residual gold in these samples is associated with pyrite, arsenopyrite or other difficult to leach sulphide minerals for both the CAP and ODM samples. The amount of the residual gold recovered by the aqua regia leach was estimated to be between 62% and 92%. Little to no gold was readily recoverable using intensive cyanide leaching, with four (4) of the six (6) samples having gold pregnant leach solution (PLS) tenors below the detection level and the other two (2) being at the detection level. Higher percentages of the residual gold were recovered using the HCl leach followed by intensive cyanide leach, with approximately 8% to 24% of the residual gold being leached using this method. Three (3) of the six (6) samples had final residual gold below detection limit (0.02 g/t) while the other three (3) samples were measured at 0.02 g/t.

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13.9

Cyanide Destruction Testwork

The SO₂/air cyanide destruction process was investigated on two (2) composites: Initial Pit and Remaining-Life-of-Mine. One (1) large bulk cyanide destruction and three (3) continuous tests were conducted for each composite.

The cyanide testwork results are presented in Table 13-29.

Table 13-29: Cyanide Destruction Testwork Results


		Pulp Density		Retention Time		Solution Phase												Reagent Addition
		Pulp Density		Retention Time		pH		CN_T		CN_WAD				Cu		Fe		g/g CN_WAD
Sample		%		Min				mg/L		mg/L¹		mg/L²		mg/L		mg/L		SO₂		Lime		Cu
Initial Pit	Feed		—		—		10.7		152		117		—		9.4		1.8		—		—		—
	Batch
	CND 3		50		90		8.6		—		—		<0.1		—		—		7.52		3.48		0.13
	Continuous
	CND 3-1		50		75		8.6		3.1		0.19		0.40		0.08		0.10		5.33		3.33		0.12
	CND 3-2		50		81		8.6		4.2		0.49		0.67		0.47		0.43		5.28		2.57		0.00
	CND 3-3		50		80		8.6		5.2		0.12		0.12		0.73		0.58		4.66		1.89		0.00
Remaining Life of Mine	Feed		—		—		11.1		128		123		—		11.0		—		—		—		—
	Batch
	CND 4		50		180		8.5		—		—		0.4		—		—		12.7		14.9		0.24
	Continuous
	CND 4-1		50		88		8.5		3.5		<0.1		0.38		0.07		0.10		4.46		4.47		0.23
	CND 4-2		50		85		8.5		3.9		<0.1		0.25		<0.05		0.13		4.17		6.71		0.25
	CND 4-3		50		99		8.5		5.8		0.13		0.29		0.10		0.52		4.24		1.79		0.00

¹	By analytical assay.

²	By picric acid.

The results showed that this process is effective at lowering the weak acid dissociable cyanide (CN_WAD) levels to well below 5 ppm. The average reagent consumptions were 4.7, 3.5 and 0.1 g/g CN_WAD for SO₂, lime and copper, respectively. The reagent consumptions are in agreement with standard industrial practices.

13.10

Carbon-in-Pulp Modelling

Carbon-in-Pulp (“CIP”) modelling work was performed to validate the design of the CIP circuit. This technique is usually used for modelling of conventional CIP circuits; however, it has been modified to model the kinetics of a carousel-style pump cell CIP circuit. Only gold is modelled by SGS.

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The Initial Pit and Remaining-Life-of-Mine composites were used for the CIP modeling testwork.

The isotherms from the testwork are presented in Figure 13-17.

LOGO

Figure 13-17: CIP Modeling Isotherms

Using the isotherms, it is possible to model the kinetics of gold adsorption onto carbon in CIP. The adsorption kinetics are modelled using a kK value, which is the product of the model output kinetic constant (k) and the model output equilibrium constant (K). The kK values from the testwork were 69 and 79 for the Initial Pit and Remaining-Life-of-Mine composites, respectively. These values are slightly lower than the reference value of 100 which is usually used as a cut-off from slow to fast kinetics.

Modeling was performed by SGS to investigate the effect of number of CIP tanks, frequency of carbon movement and size of CIP tanks on adsorption efficiency. The simulations yielded solution losses of between 0.009 to 0.035 mg/L, depending on the configuration. The results indicated that a 7- or 8-tank configuration is required to achieve high gold adsorption efficiency and that the ability to transfer carbon every day is beneficial. Based on these results, the CIP circuit was designed to have seven (7) tanks and the stripping circuit sized to be able to strip and regenerate 100% of the carbon every day.

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It should be noted that the numbers from this model are considered to be a worst case scenario and conservative. It is not recommended to use these numbers for the financial analysis of a project, but rather for sizing the equipment. It was recommended that a constant residual solution gold tenor of 0.007-0.008 mg/L be used for the financial analysis based on the kinetics observed from the two (2) composites.

13.11

Thickener Sizing Testwork

13.11.1

Flocculant Screening

Flocculant screening was performed by a flocculant supplier, prior to performing sedimentation rate testwork. This was to ensure that all of the sedimentation tests performed by three (3) thickener suppliers were done using the same basis. Pre-leach and pre-detox thickener feed material were tested using two (2) samples: an Initial Pit composite and a Remaining-Life-of-Mine composite. The target grind size for the samples was a P₈₀ of 75 µm.

Four (4) different flocculants were tested and are presented in Table 13-30.

Table 13-30: Flocculant Description


		Molecular Weight		Charge Density
Flocculant	Charge	(10⁶Dalton)		(mol %)
Flocculant 1	Anionic		~13-16		5	%
Flocculant 2	Anionic		~13-16		10	%
Flocculant 3	Non-ionic		~9-11		Low
Flocculant 4	Anionic		~14-17		20	%

The testwork results indicated that the non-ionic flocculant (Flocculant 3) had the most consistent performance in terms of settling rates and overflow clarity. It was also noted that Flocculant 1 showed fairly good performance; however, Flocculants 2 and 4 did not meet requirements. Flocculant 2 had poor overflow clarity for several of the samples, while Flocculant 4 had little effect on either settling rates or overflow clarity. The results clearly indicate that the flocculant selected should have a low charge density and be non-ionic or slightly anionic. It was determined that the sedimentation testwork would be performed with the non-ionic Flocculant 3 (or equivalent).

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It was also noted that dual addition of flocculant was required for certain samples to achieve acceptable overflow clarity. Furthermore, a pH of 10.5 or higher significantly improved the settling rates and overflow clarity.

13.11.2

Sedimentation Testwork

Sedimentation testwork was performed at three (3) different supplier laboratories to size the pre-leach and pre-detox thickeners as there is often a significant degree of variation between the different laboratories and methods.

The sedimentation testwork results are presented in Table 13-31.

Table 13-31: Sedimentation Testwork Results


Sample	Description	Units	Supplier A				Supplier B				Supplier C
Sample	Description	Units	Pre- Leach		Pre- Detox		Pre- Leach		Pre- Detox		Pre- Leach	Pre- Detox
Design Feed Rate (Dry)		t/h		906		906		906		906	906	906
Initial Pit	Settling Rate	t/h/m²		0.65		0.86		0.90		0.90	0.61 – 1.05	0.61 - 1.05
	Rise Rate	m/h		<7		<7		—		—	3.4 – 5.9	3.4 - 6.0
	Flocculant Dosage	g/t		30-35		40-45		40		30	20 – 40	20 -41
	Overflow Clarity	ppm		<200		<200		<150		<150	10 -86	9 – 57
Remaining-Life-of-Mine	Settling Rate	t/h/m²		—		—		1.0		1.0	0.65 – 1.14	0.65 – 1.11
	Rise Rate	m/h		—		—		—		—	3.6 – 6.3	3.6 – 6.1
	Flocculant Dosage	g/t		—		—		25		50	19 – 40	29 – 48
	Overflow Clarity	ppm		—		—		<200		<200	50 – 145	29 - 205
Recommended Diameter		m		44		38		38		38	45	45

The results indicated that the recommended thickener diameter is between 38 and 45 m. The lowest settling rates were observed by Supplier C, while the highest were observed by Supplier B. Based on these results, it is recommended that 44 m diameter pre-leach and pre-detox thickeners be used.

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It can also be seen that the flocculant dosage requirements were fairly high. The flocculant dosage ranged from 19 to 40 g/t (with an average for the three (3) suppliers of around 32 g/t) and 20 to 48 g/t (with an average for the three (3) suppliers of around 39 g/t). These results were seen consistently across the three (3) supplier laboratories with little variation.

13.12

Rheology

Rheology testwork was performed on the Initial Pit and RLOM composites using a concentric cylinder rotational viscometer (“CCRV”). The objective of the testwork was to determine the yield stresses at different percent solids for an unsheared and sheared sample. The yield stress is the minimum force that is required to cause of fluid to flow. Using these values, it is possible to determine the Critical Solids Density (“CSD”). The CSD is the density at which an incremental increase in the solids density causes a significant increase in yield stress, or a significant decrease in flowability of the slurry. The CSD is also considered to be a prediction of the maximum underflow solids density that can be achieved by a commercial thickener for the material and is considered to be the optimum thickener underflow percent solids.

The rheology results are presented in Figure 13-18 and Figure 13-19.

LOGO

Figure 13-18: Initial Pit Composite Yield Stress vs. Solids Density

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LOGO

Figure 13-19: Remaining-Life-of-Mine Composite Yield Stress vs. Solids Density

It can be seen that the CSD was 62.2% w/w and 63.5% w/w for the Initial Pit and RLOM composites, respectively. These values are in line with the design of the pre-leach and pre-detox thickener underflow densities of 62% and 60%, respectively.

13.13

Linear Screen Sizing Testwork

Testwork was performed to determine the sizing of the trash and safety linear screens. This was done at a supplier laboratory by determining the flux rate (m³/m² /h) at different feed percent solids.

The results are presented in Table 13-32.

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Table 13-32: Linear Screen Sizing Testwork Results


Description	Units	Trash Screen					Safety Screen
Feed Volume	m³/h		3,172					1,290
% Solids	% w/w		25	25	25	25		50
Screen Aperture	µm		600	600	700	700		600
Draining Rate	m³/m ²/h		95	95	109	109		51
Required Screen Area	m²		33.4	33.4	29.1	29.1		25.3
Design Safety Factor	%		20	20	20	15		20
Required Screen Area	m²		40.1	40.1	34.9	33.5		30.4
Number of Screens			1	2	1	1		1
Screen Size	m²		40	20	40	40		32

The testwork indicated a flux rate to the trash screen of 95 and 109 m³/m²/h with screen apertures of 600 and 700 um, respectively. The safety screen indicated a flux rate of 51 m³/m²/h for a 600 um screen aperture. Based on these results, two (2) 20 m² trash screens and one (1) 32 m² safety screen were selected with screen apertures of 600 um.

13.14

Environmental Testwork

AMEC Environment and Infrastructure is conducting environmental geochemical characterization of selected samples representative of the mine rock and overburden in the vicinity of the proposed Rainy River Gold Project open pit and tailings deposition areas. Table 13-7 summarizes the geochemical testwork. To date, testing has been carried out on three (3) simulated tailings materials, and a total of 659 deposit-wide mine rock samples, of which 366 represent in-pit non-ore mine rock.

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Table 13-33: Summary of Geochemical Environmental Testing


Analytical Procedure	Mine Rock	Overburden	Tailings
Static Testing
Acid-base Accounting*	x	x	x
Net Acid Generation test	x	x	x
Metals Content (aqua-regia ICP)	x	x	x
Short-term Leach testing (SFE**)	x	x	x
Mineralogy (X-Ray Diffraction)	x		x
Whole Rock (X-ray Fluorescence)			x
Kinetic Testing
Humidity Cell	x		x
Field Cell	x

*	Sobek with siderite correction modified Sobek, total inorganic carbon and sulphur speciation.

**	SFE = 24 hour shake flask extraction at 3:1 deionized water to solid.

Geochemical studies to-date on mine rock indicate that approximately half the samples may have the potential to produce acid rock drainage. A block model is currently under development to refine the estimated tonnage of potentially acid generating rock. Results of the tailings analyses indicate a risk for acidic drainage in the future if not appropriately managed. Generally, metal contents in waste materials are typical for their rock types and the risk for metal leaching under neutral conditions appears to be low. Humidity cell analysis is currently being undertaken on mine rock and tailings to evaluate the long-term metal leaching characteristics of these materials.

13.15

Gold and Silver Recovery Curves

Gold and silver recovery curves were developed for the Financial Analysis based on the leaching and gravity recovery results presented in Sections 13.6 and 13.8.6 and the proposed whole rock leach flowsheet, discussed in Section 17. The recovery curves were developed using the following format:

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Where:

•

Rec_Au is the overall gold recovery;

•

Rec_Ag is the overall silver recovery;

•

x_Au is the gold head grade;

•

x_Ag is the silver head grade;

•

A₍_{Au Res)}is the gold residue;

•

A₍_{Ag Res)}is the silver residue;

•

B₍_{Au Gravity} _Rec₎is the gold gravity recovery;

•

B₍_{Ag Gravity} _Rec₎is the silver gravity recovery;

•

CIP_eff-Au is the CIP adsorption efficiency for gold;

•

CIP_eff-Ag is the CIP adsorption efficiency for silver;

•

EW_eff is the CIP-stripping solution electrowinning (EW) recovery; and

•

IL_eff is the intensive cyanidation and dedicated electrowinning efficiency.

The residue and gravity recovery curves were developed based on grade-recovery variability testwork and divided into CAP and non-CAP Zones (non-CAP Zones include ODM, Z-433, NZ and CAP). These curves are presented below:

Table 13-34: Residue and Gravity Recovery Curves


Non-CAP Zones		CAP Zone
A_{(Au Res)} = Cx^d = 0.0937. x_Au^0.4223		A_{(Au Res)} = Cx^d = 0.2497.x_Au^1.015

B_{(Au Gravity Rec)} = Ex + F = (11.15 . x_Au + 17.3)		B_{(Au Gravity Rec)} = Ex + F = (8.58.x_Au + 2.94)

A_{(Ag Res)} = Gx² + Hx + I = 0.01 .x_Ag ² + 0.29x_Ag		A_{(Ag Res)} = Gx² + Hx + I = 0.036.x_Ag² + 0.244x_Ag

B_{(Ag Gravity Rec)} = Ex + F = 10.0		B_{(Ag Gravity Rec)} = Ex + F = 3.6

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The CIP_eff-Au was calculated using a fixed discharge solution gold tenor of 0.007 mg/L based on testwork performed by SGS. Testwork for silver modeling was not performed, and therefore the model was based on discussions with CIP suppliers. The silver adsorption efficiency was estimated to be 96.6%. EW_eff and IL_eff values were assumed to be 100% since all residual solids and solutions from these circuits are recycled to the process.

The expected gold and silver recoveries as a function of head grade (gold or silver, respectively) are presented in Table 13-35.

Table 13-35: Gold and Silver Recoveries vs. Head Grade


	Gold Recovery				Silver Recovery
Head Grade (g/t)	Non-CAP		CAP		Non-CAP		CAP
0.2		72.8		72.1		68.2		72.4
0.4		82.3		73.6		67.9		71.7
0.6		86.2		74.1		67.6		71.0
0.8		88.5		74.2		67.4		70.3
1		89.9		74.3		67.1		69.6
1.2		91.0		74.4		66.8		68.9
1.4		91.8		74.4		66.5		68.2
1.6		92.4		74.4		66.2		67.5
1.8		92.9		74.4		65.9		66.8
2		93.4		74.4		65.6		66.1
2.5		94.2		74.4		64.9		64.3
3.0		94.8		74.4		64.2		62.6
3.5		95.3		74.4		63.5		60.8
4.0		95.6		74.3		62.8		59.1
4.5		95.9		74.3		62.1		57.3
5.0		96.2		74.3		61.4		55.6

As previously noted, it can be seen that the gold recovery for the CAP Zone is considerably lower than the non-CAP Zones. The annual ratio of Non-CAP and CAP material will have to be an input for the financial analysis to calculate the average gold recovery by year. The silver recovery was identical for both the non-CAP and CAP Zones and is independent of the head grade.

13.16

Testwork Interpretation

The results from the SGS testwork are the basis for the mineral reserve estimate and Feasibility Study. Based on a trade-off study, it was determined that the whole rock leaching option with

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gravity separation was the most economical alternative and was therefore used as the basis for the Feasibility Study. The main reason for this selection was the significant amount of energy associated with regrinding the flotation concentrate and the high cyanide consumption in the flotation concentrate leaching, in addition to risk associated with ultrafine grinding of this material. All subsequent testwork was based on cyanide leaching of the gravity tailings.

The grinding tests indicate that significant variation exists in the mineral hardness in the ODM Zone, and that the overall deposit is considerably harder than previously indicated in the December 2011 PEA. The design Axb value was modified from 34.0 to 24.2, resulting in an increase of close to 50% in terms of power requirements for the SAG mill. The extensive grinding testwork campaign has allowed for definition of the overall hardness of each zone and indicated that there are several portions of the deposit that will have high energy requirements and this will be reflected in the design of the process plant. The strong correlation between the four (4) methods used to size the grind circuit provides a good level of confidence in the sizing of the SAG and ball mill.

The process is expected to yield an overall gold recovery of approximately 90 to 91% and a silver recovery of around 66 to 67% over the life-of-mine without considering solution losses. When considering solution losses, the gold recovery decreases by approximately 0.4% while the silver recovery drops to approximately 64%. The grind size chosen for this study was 75 µm, based on a cost versus revenue study performed by BBA. The gold recovery varies significantly throughout the deposit and the CAP Zone has considerably lower gold recoveries than the other zones. As per the mine plan schedule, CAP Zone material, when mined, is placed in the low grade ore stockpile and therefore only treated towards the end of the mine life. No CAP Zone material is processed in Years 1-10, resulting in an elevated recovery for those years.

Average gold and silver recoveries are based on the mine plan and shown in Table 13-36.

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Table 13-36: Average Yearly Gold and Silver Recoveries


	Gold				Silver
Years	Au Head Grade (g/t)		Au Recovery (%)		Ag Head Grade (g/t)		Ag Recovery (%)
1		1.19		90.9		3.13		64.0
2		1.31		91.5		2.99		64.2
3		1.36		91.6		3.85		63.0
4		1.76		92.9		2.11		65.5
5		1.68		92.6		2.43		65.0
6		1.54		92.2		2.73		64.6
7		1.32		91.5		3.35		63.7
8		1.31		91.4		5.03		61.3
9		1.35		91.6		3.85		63.0
10		1.58		92.4		2.41		65.1
11		0.64		86.8		2.31		65.2
12		0.53		85.2		2.27		65.3
13		0.49		84.4		1.50		66.4
14		0.32		79.8		1.92		65.7
15		0.30		78.8		2.10		65.5
16		0.52		74.4		2.28		64.9
LOM		1.08		90.4		2.76		64.1

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14.

MINERAL RESOURCE ESTIMATION

14.1

Introduction

In May 2012, SRK was commissioned by Rainy River Resources Ltd. to prepare an updated mineral resource model for the Rainy River Gold Project to consider new drilling data available up to July 10, 2012. The resource estimation work was completed in Toronto, by Dorota El-Rassi, P.Eng (PEO #100012348) and Glen Cole, P.Geo. (APGO #1416), both “independent qualified persons” as this term is defined in National Instrument 43-101.

The Mineral Resource Statement presented herein represents the eighth mineral resource evaluation prepared for the Rainy River Gold Project since 2003.

The Rainy River Gold Project contains volcanic hosted gold-rich polymetallic sulphide mineralization of hydrothermal origin that is crosscut by a small zone of magmatic copper-nickel sulphide mineralization enriched in platinum group metals. The Mineral Resource Statement is reported on the basis of gold content only, although locally significant silver is also present. The consolidated Mineral Resource Statement for the Rainy River Gold Project reports gold and silver grades only. The mineral resources for the copper-nickel sulphide mineralization enriched and silver enriched zones are reported separately because they contain more substantial base metal and silver mineralization respectively.

This section describes the resource estimation methodology used by SRK and summarizes the key assumptions and parameters used to prepare the eighth Mineral Resource Statement prepared for the Rainy River Gold Project.

In the opinion of SRK, the resource evaluation reported herein is a reasonable representation of the mineral resources found in the Rainy River Gold Project at the current level of sampling. The mineral resources have been estimated in conformity with generally accepted CIM “Estimation of Mineral Resource and Mineral Reserves Best Practices” guidelines and are reported in accordance with Canadian Securities Administrators’ NI 43-101. Mineral resources are not mineral reserves and do not have demonstrated economic viability. There is no certainty that all or any part of the mineral resource will be converted into a mineral reserve.

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14.2

Resource Estimation Procedures

The evaluation of mineral resources for the Rainy River Gold Project involved the following procedures:

•

Database compilation and verification;

•

Construction of wireframe models for major lithological units, using stratigraphy, structural trends and an array of appropriate geochemical indices;

•

Definition of geostatistical resource domains;

•

Data conditioning (compositing and capping) for geostatistical analysis and variography;

•

Selection of estimation strategy and estimation parameters;

•

Block modeling and grade interpolation;

•

Validation, classification and tabulation;

•

Assessment of “reasonable prospects for economic extraction” and selection of reporting cut-off grades; and

•

Preparation of Mineral Resource Statement.

14.3

Resource Database

Exploration data used to evaluate the mineral resources for the Rainy River Gold Project were provided by Rainy River as a set of Microsoft Excel files containing drilling information (drill collars, surveys, assays and lithological logging) for 1,435 core boreholes (662,849 m) drilled by Rainy River and Nuinsco, the previous Project operator. The database includes 237 boreholes (95,760 m) drilled in 2012, 388 boreholes (188,588 m) drilled by Rainy River in 2011 and early 2012, 165 boreholes (84,133 m) drilled by Rainy River in 2010, 446 boreholes (245,016 m) drilled by Rainy River between 2005 and 2009, and 199 boreholes (49,351 m) drilled by Nuinsco between 1994 and 2004.

All exploration information is located using the local UTM grid (NAD 83 datum, Zone 15). Resource modelling was conducted in this UTM coordinate space.

A topography surface was also supplied to SRK by Rainy River in DXF format. Upon receipt of the digital drilling data, SRK performed the following validation steps:

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•

Checked minimum and maximum values for each quality value field and confirmed/edited those outside of expected ranges;

•

Checked for inconsistencies in lithological unit terminology and/or gaps in the lithological table; and

•

Checked for gaps, overlaps, and out of sequence intervals for both assays and lithology tables.

SRK has previously reviewed analytical quality control data produced by Rainy River for the periods prior to December 2011 (documented in previous technical reports). For this Study, SRK reviewed the analytical quality control data produced between December 2011 and June 2012. The analytical quality control data produced during this period is summarized on bias charts and precision plots presented in Appendix D.

Each interval in the assay table was assigned a new rock code value based on the location of the interval mid-point relative to the modelled gold mineralization. The coded assay data were extracted for statistical analysis.

The database used to estimate the mineral resources was audited by SRK. SRK is of the opinion that the current drilling information is sufficiently reliable to interpret the outlines of the gold mineralization with reasonable confidence, and that the assay data are sufficiently reliable to support mineral resource estimation.

Leapfrog™ software was used to guide geology and mineralization modelling. Gemcom GEMS™ software was used to construct the geological solids, prepare assay data for geostatistical analysis, construct the block model, estimate metal grades, and tabulate mineral resources. The Geostatistical Software Library™ (GSLib) family of software and GEMS were used for geostatistical analysis and variography. Conceptual pit optimization work to test the “reasonable prospects” for economic extraction was completed with MineSight software by BBA.

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14.4

Solid Body Modelling

14.4.1

Introduction

SRK and Rainy River constructed a series of 3D wireframes to constrain the extent of the gold mineralization, considering structural features, lithology, alteration, geochemical indices, as well as grade trends. Rainy River generated gold mineralization wireframes for Zone 433 as well as the CAP and HS Zones, whereas SRK generated gold mineralization wireframes for all the other Zones. From these wireframes, resource domains were constructed and used as hard boundaries to constrain grade estimation. The resource domains were updated by SRK to consider the new drilling information (Figure 14-1). The geological interpretation and resource domains have not changed significantly relative to the previous resource model with the exception of the New Zone, which was amalgamated with the HS Zone.

The Rainy River Gold Project was subdivided into 12 separate resource domains: six (6) main Zones (ODM/17, 34, 433, HS, CAP, Western), the Silver Zone; and five (5) isolated pockets outside the main Zones (Table 14-1). The 3D shapes for the six (6) main Zones are illustrated in Figure 14-2. The ODM/17 and 433 Zones were subdivided into three (3) grade subdomains (Low, Medium and High grade).

Table 14-1: Rock Codes in the Rainy River Gold Project Block Model


Zone	Domain	Domain Code
ODM/17	Low Grade (below 0.5 g/t gold)		101 to 102
	Medium Grade (between 0.5 and 0.9 g/t gold)		110 to 114
	High Grade (above 0.9 g/t gold)		120 to 123
Zone 34			200
Zone 433	Low Grade (below 0.5 g/t gold)		300
	Medium Grade (between 0.5 and 0.9 g/t gold)		310
	High Grade (above 0.9 g/t gold)		320
HS			400
CAP			500
Mineralization outside Main Zones			601 to 605
Western			800
Silver			901 to 904

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Figure 14-1: Location of New Boreholes Drilled During the Period March to December 2011,

Relative to Previous Drilling

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Figure 14-2: Isometric View of the Rainy River Mineralization Wireframes Modelled by SRK

with Borehole Data (View looking towards the west showing 2011 drilling only)

The methodology adopted for modelling the main Zones of gold mineralization are summarized in the following paragraphs.

14.1.1

The ODM/17 Zone

The ODM/17 Zone (Figure 14-1 and Figure 14-2) is interpreted as a generally east-west trending, south-west plunging Zone of mineralization. Numerous north-north-east striking faults crosscut the Zone. Small offsets along the Beaver Pond Fault were modelled. However, at the scale of the Zone, none of these faults were used as hard boundaries. Numerous alteration indices, as well as gold grade shells, suggest a stacked pattern of slightly oblique Zones that resemble tight folds. The general outline of the ODM/17 Zone was based on the broad extent of a sericite index (cationic based) larger than 0.7. The outlines initially were guided by a 3D model of the sericite index and a 0.2 g/t gold grade shell.

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The hanging wall of the Zone coincides with the top of a fragmental volcaniclastic unit that hosts much of the ODM/17 Zone. This rock package is separated from mafic volcanic and intermediate to felsic volcanic rock to the south by a curved but generally east-west trending magnetic lineament. This lineament was modelled and used as the hanging wall boundary of the ODM/17 Zone. This contact becomes cryptic to the east, but was extended parallel to the magnetic lineament. The ODM/17 domain was defined on inclined sections oriented perpendicular to the plunge (azimuth 233 degrees plunge of 47 degrees). The overall ODM/17 Zone was subdivided into three (3) grade subdomains based on the following divisions:


• High Grade:		Greater than 0.9 g/t gold;

• Medium Grade:		Between 0.5 and 0.9 g/t gold, and

• Low Grade:		Below 0.5 g/t gold.

The ODM/17 Zone was therefore subdivided into three (3) resource domains that were considered for resource estimation. The geometry of the Medium and High grade subdomains is either parallel to the south-dipping footwall of the overall domain or slightly oblique to it. These are consistent with the geometry of high strain zones bounding the subdomains and the foliation orientation within them.

14.1.2

The 433, HS and New Zones

The 433 and HS Zones (Figure 14-1 and Figure 14-2) form two (2) of several gold occurrences in the footwall of the ODM/17 Zone, hosted by massive and fragmental felsic to intermediate rock. The boundaries of these Zones are not as well defined as for the ODM/17 Zone, but the gold mineralization plunge is similar. Accordingly, the boundaries for the 433 and HS Zones were modelled on the same inclined sections. The sericite index does not clearly define these Zones. In the case of the 433 Zone, chalcopyrite is associated with gold mineralization and is a good indicator of the boundaries of that Zone.

The boundaries were modelled using a copper-to-zinc ratio of 0.8. Similarly to the ODM/17 Zone, the overall 433 Zone was also subdivided into three (3) grade subdomains based on the following divisions:

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• High Grade:		Greater than 0.9 g/t gold;

• Medium Grade:		Between 0.5 and 0.9 g/t gold, and

• Low Grade:		Below 0.5 g/t gold.

The 433 Zone was therefore subdivided into three (3) resource domains that were considered for resource estimation.

No geochemical or lithological criteria successfully outline the HS Zone. The HS Zone was defined with consideration of the lithogical model and by using the interpreted extent of a 0.2 g/t gold threshold (based on 3 m composites) and guided by 0.2 g/t gold Leapfrog™ shells.

The additional infill drilling information suggests that the former New Zone is part of the HS Zone. Accordingly the two (2) were combined, effectively increasing the size of the HS Zone considerably compared to that in earlier models.

The CAP Zone

The CAP Zone occurs in the hanging wall of the ODM/17 Zone (Figure 14-1 and Figure 14-2) within the upper, predominantly mafic volcanic sequence. On the surface, the Zone is associated with a number of quartz-carbonate vein sets and south-dipping shear zones. The latter post-date the foliation and dip both more steeply and more shallowly than the south-dipping foliation. The orientation of the quartz-carbonate veins is highly variable. North-east to north-west striking sulphide veinlets anastomose across several surface outcrops. In core individual high-grade gold intersections are associated with increased sulphide mineralization (particularly chalcopyrite) within and adjacent to shear Zone hosted quartz-carbonate veins.

Low-grade gold mineralization in intermediate rocks within the CAP Zone is similar to the ODM/17 Zone, with a noticeably shallower plunge to the south-west. On north-south vertical sections, high-grade gold intersections are aligned along south-dipping planes. In plan view, high-grade gold intersections show continuity along a west-north-west strike. Low-grade mineralization shows good continuity when observed in cross-sections oriented perpendicular to the slightly shallower plunge. The CAP Zone domain was modelled on vertical sections based on a 0.2 g/t gold threshold guided by this preferred geometry.

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Western Zone

Gold mineralization appears sporadic in the Western Zone of the Rainy River Gold Project, but can be subdivided into at least two (2) stages of mineralization:

•

Early (low- to moderate-grade) gold mineralization associated with sulphide (pyrite-sphalerite-chalcopyrite-galena) stringers and veins and disseminated pyrite in quartz-phyric volcaniclastic rocks and conglomerate; and

•

Late (high-grade) gold mineralization associated with quartz-carbonate-pyrite-gold veins and veinlets and rarely as native gold veins.

This hybrid mineralization consists of an early gold-rich volcanogenic sulphide mineralization overprinted by shear-hosted mesothermal gold mineralization. Gold mineralization is commonly associated with increased sericite and chlorite alteration. Mineralization also appears to have a strong association with strain. Increased strain, characterized by kink folds, boudinage, and strong fabric development, is commonly associated with an increased gold grade. However, at very high strain, mylonitic textures appear and the gold grade is reduced. The Western Zone can be interpreted as a north-west extension of the ODM/17 Zone. The domain was defined on vertical sections guided by 0.2 g/t gold Leapfrog™ shells. At present, gold mineralization in the Western Zone appears erratic and discontinuous; infill drilling would be required to improve grade continuity.

The 34 Zone

The 34 Zone was modelled by Rainy River and modified by SRK using drilling logging data and Leapfrog™. The model was used to constrain mineral resource estimation.

Silver Zone

The Silver Zone occurs in the footwall of the ODM/17 Zone in dacitic tuff and breccias immediately adjacent to a high strain Zone located at the northern contact of the ODM/17 Zone. The Zone plunges to the south-west in similar orientation to the ODM/17 Zone, and is associated with centimetre-scale sulphide bearing quartz veinlets that typically contains dendritic native silver inclusions. The Silver Zone domain was outlined by Rainy River by using a 19 g/t silver cut-off

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grade (3.0 m composites; less than 4.0 m waste), on inclined cross-sections perpendicular to the plunge of the silver mineralization.

14.5

Compositing

Most of the core assay samples were taken at 1.5 m intervals. A histogram of the raw assay lengths inside the mineralized envelopes is provided in Figure 14-3. For geostatistical analysis, variography and grade estimation, raw assay data were composited to equal 1.5 m lengths. Compositing was completed from entry point of the wireframe, down the hole. Composite residuals that were shorter than 10% of composite length were removed from the data set.

LOGO

Figure 14-3: Histogram Distribution of Raw Sample Lengths

14.6

Evaluation of Outliers

SRK constructed cumulative probability curves for the gold and silver composites within each modelled domain (shown in Figure 14-4). Considering the nature of the statistical distributions of gold assay, SRK is of the opinion that it is necessary to cap high-grade values to limit their influence during grade estimation. The impact of capping was analyzed and capping levels were

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adjusted for each resource domain (and subdomains therein) and each metal separately. Capping was applied to the composites. Capping levels for gold and silver are summarized in Table 14-2.

Table 14-2: Summary of Metal Capping Levels Applied to Each Resource Domain


	Gold							Silver
Resource Domain	Cap (g/t)		Percentile			N capped		Cap (g/t)		Percentile			N capped
100		20		99.97	%		19		100		99.97	%		20
MG 110 to 114		60		99.93	%		10		120		99.91	%		13
HG 120 to 123		115		99.86	%		12		80		99.93	%		6
200		3		98.71	%		6		30		99.35	%		3
300		25		99.97	%		3		35		99.97	%		4
310		30		99.75	%		7		40		99.78	%		6
320		70		99.29	%		7		15		99.19	%		8
400		25		99.94	%		6		40		99.92	%		8
500		15		99.95	%		6		70		99.92	%		9
601		30		99.99	%		8		30		99.87	%		67
602		9		99.98	%		5		50		99.98	%		5
603		20		99.99	%		3		150		99.99	%		4
604		10		99.99	%		9		70		99.98	%		14
605		5		99.98	%		7		30		99.99	%		4
800		20		99.26	%		8		55		99.63	%		4
901		0.8		95.42	%		7		250		97.39	%		4
902		9		96.47	%		3		80		94.12	%		5
903		4		97.79	%		3		70		97.79	%		3
904		4		99.52	%		2		115		99.29	%		3

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Figure 14-4: Cumulative Frequency Plot for Gold Composites

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14.7

Statistical Analysis and Variography

14.7.1

Statistical Analysis

The basic statistics for the composite and capped composite data within the 11 resource domains for gold and silver are summarized in Table 14-3 to Table 14-6.

The basic statistics for the raw, composited, and capped composited data of the various metals in Domain 200 (34 Zone) are summarized in Table 14-7.

SRK also undertook a domainal statistical analysis of sulphur and calcium extracted from the ICP database. Sulphur and calcium were also interpolated into the block model for use in waste rock characterization and waste management disposal.

The basic statistics for the composite and capped calcium and sulphur composites within the eleven resource domains are summarized in Table 14-8 to Table 14-11.

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Table 14-3: Basic Statistics for Gold Composites for All Resource Domains


Zone	Domain	Subdomain		Count		Minimum		Maximum		Mean		Standard Deviation		Sample Variance

	Low		100		74,758		0.00		448.01		0.20		1.77		3.15
			110		3,337		0.00		61.78		0.72		2.22		4.94
			111		3,857		0.00		1219.63		1.08		19.82		392.69
			112		3,490		0.00		110.14		0.87		3.31		10.92
	Medium		113		2,007		0.00		60.76		0.73		1.93		3.72
			114		838		0.03		1844.65		3.16		63.83		4073.83
ODM/17			115		722		0.00		53.40		1.22		3.00		8.99
			mg_110_114		14,251		0.00		1844.65		1.02		18.73		350.66
			120		2,042		0.00		195.28		1.82		6.66		44.33
			121		3,047		0.00		301.73		2.10		7.65		58.48
	High		122		3,191		0.00		213.72		2.74		9.57		91.61
			123		544		0.00		462.03		2.92		21.20		449.48
			hg_120_123		8,824		0.00		462.03		2.32		9.56		91.39
Zone 34	Low		200		465		0.00		26.57		0.35		1.64		2.70
	Low		300		11,553		0.00		60.19		0.27		1.04		1.08
433	Medium		310		2,774		0.00		629.23		1.10		12.30		151.31
	High		320		986		0.00		2275.98		5.43		75.44		5691.65
HS			400		10,276		0.00		249.45		0.48		3.04		9.24
CAP			500		11,477		0.00		64.77		0.39		1.13		1.27
			601		53,369		0.00		108.06		0.08		0.76		0.58
			602		30,264		0.00		87.72		0.09		0.67		0.45
600 Series			603		52,241		0.00		104.98		0.08		0.58		0.34
			604		83,114		0.00		58.55		0.05		0.31		0.10
			605		38,934		0.00		43.93		0.03		0.33		0.11
Western			800		1,078		0.00		324.98		1.15		11.15		124.29
			901		153		0.00		4.04		0.23		0.56		0.31
Silver Zones			902		85		0.11		1039.62		18.75		117.11		13714.41
			903		136		0.02		19.28		0.95		2.02		4.10
			904		420		0.00		6.35		0.44		0.70		0.49

14-14

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

Table 14-4: Base Statistics for Capped Gold Composites for All Resource Domains


Zone	Domain	Subdomain		Count		Minimum		Maximum		Mean		Standard Deviation		Sample Variance
	Low		100		74,758		0.00		20.00		0.19		0.55		0.31
			110		3,337		0.00		60.00		0.72		2.20		4.86
			111		3,857		0.00		60.00		0.76		2.53		6.40
			112		3,490		0.00		60.00		0.86		2.83		8.04
	Medium		113		2,007		0.00		60.00		0.73		1.92		3.67
			114		838		0.03		60.00		0.99		4.00		16.01
ODM/17			115		722		0.00		53.40		1.22		3.00		8.99
			mg_110_114		14,251		0.00		60.00		0.81		2.60		6.76
			120		2,042		0.00		115.00		1.75		5.08		25.80
			121		3,047		0.00		115.00		2.02		5.48		30.01
	High		122		3,191		0.00		115.00		2.64		7.87		61.89
			123		544		0.00		115.00		2.18		7.72		59.61
			hg_120_123		8,824		0.00		115.00		2.19		6.52		42.50
Zone 34	Low		200		465		0.00		3.00		0.24		0.49		0.24
	Low		300		11,553		0.00		25.00		0.26		0.79		0.63
433	Medium		310		2,774		0.00		30.00		0.83		2.05		4.22
	High		320		986		0.00		70.00		2.34		7.36		54.14
HS			400		10,276		0.00		25.00		0.45		1.11		1.23
CAP			500		11,477		0.00		15.00		0.38		0.77		0.59
			601		53,369		0.00		30.00		0.08		0.48		0.23
			602		30,264		0.00		9.00		0.09		0.24		0.06
600 Series			603		52,241		0.00		20.00		0.08		0.32		0.10
			604		83,114		0.00		10.00		0.04		0.18		0.03
			605		38,934		0.00		5.00		0.03		0.11		0.01
Western			800		1,078		0.00		20.00		0.67		2.01		4.05
			901		153		0.00		0.80		0.16		0.20		0.04
Silver Zones			902		85		0.11		9.00		1.99		2.28		5.20
			903		136		0.02		5.27		0.78		0.96		0.93
			904		420		0.00		6.18		0.43		0.66		0.43

14-15

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

Table 14-5: Basic Statistics for Silver Composites for All Resource Domains


Zone	Domain	Subdomain		Count		Minimum		Maximum		Mean		Standard Deviation		Sample Variance
	Low		100		74,758		0.00		733.61		1.31		5.87		34.50
			110		3,337		0.00		230.91		3.06		8.85		78.24
			111		3,857		0.00		99.87		1.29		2.98		8.88
			112		3,490		0.00		42.00		1.28		2.34		5.45
	Medium		113		2,007		0.00		99.98		2.54		5.61		31.53
			114		838		0.00		90.39		4.55		7.73		59.79
ODM/17			115		722		0.00		537.98		9.39		31.14		969.39
			mg_110_114		14,251		0.00		537.98		2.48		9.09		82.54
			120		2,042		0.00		221.02		4.55		10.53		110.82
			121		3,047		0.00		114.00		2.11		4.35		18.94
	High		122		3,191		0.00		66.57		2.21		3.96		15.71
			123		544		0.00		121.84		2.59		6.21		38.60
			hg_120_123		8,824		0.00		221.02		2.74		6.42		41.23
Zone 34	Low		200		465		0.00		40.19		1.95		4.94		24.38
	Low		300		11,553		0.00		98.00		0.53		1.69		2.85
433	Medium		310		2,774		0.00		99.62		0.79		3.34		11.15
	High		320		986		0.00		240.25		1.25		8.35		69.71
HS			400		10,276		0.00		672.56		0.97		7.04		49.52
CAP			500		11,477		0.00		440.33		2.04		6.26		39.25
			601		53,369		0.00		57.63		0.45		1.59		2.54
			602		30,264		0.00		90.58		0.43		1.64		2.70
600 Series			603		52,241		0.00		476.44		0.59		3.47		12.04
			604		83,114		0.00		1110.23		0.37		4.41		19.49
			605		38,934		0.00		151.07		0.36		1.36		1.85
Western			800		1,078		0.00		125.35		1.64		7.03		49.36
			901		153		0.00		334.11		59.34		67.10		4502.40
Silver Zones			902		85		0.23		300.52		24.75		39.28		1543.07
			903		136		0.00		131.98		16.33		19.35		374.29
			904		420		0.00		204.05		15.86		22.65		512.83

14-16

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

Table 14-6: Base Statistics for Capped Silver Composites for All Resource Domains


Zone	Domain	Subdomain		Count		Minimum		Maximum		Mean		Standard Deviation		Sample Variance
	Low		100		74,758		0.00		100.00		1.27		3.36		11.30
			110		3,337		0.00		120.00		2.98		7.21		51.94
			111		3,857		0.00		99.87		1.29		2.98		8.88
			112		3,490		0.00		42.00		1.28		2.34		5.45
	Medium		113		2,007		0.00		99.98		2.54		5.61		31.53
			114		838		0.00		90.39		4.55		7.73		59.79
ODM/17			115		722		0.00		120.00		7.89		16.72		279.55
			mg_110_114		14,251		0.00		120.00		2.38		6.36		40.51
			120		2,042		0.00		80.00		4.34		8.15		66.50
			121		3,047		0.00		80.00		2.09		3.91		15.27
	High		122		3,191		0.00		66.57		2.21		3.96		15.71
			123		544		0.00		80.00		2.51		4.84		23.44
			hg_120_123		8,824		0.00		80.00		2.68		5.35		28.62
Zone 34	Low		200		465		0.00		30.00		1.91		4.70		22.05
	Low		300		11,553		0.00		35.00		0.52		1.33		1.78
433	Medium		310		2,774		0.00		40.00		0.75		2.46		6.04
	High		320		986		0.00		15.00		0.87		1.91		3.65
HS			400		10,276		0.00		40.00		0.90		2.04		4.14
CAP			500		11,477		0.00		70.00		1.97		3.75		14.08
			601		53,369		0.00		30.00		0.45		1.49		2.22
			602		30,264		0.00		50.00		0.42		1.46		2.13
600 Series			603		52,241		0.00		150.00		0.58		2.49		6.22
			604		83,114		0.00		70.00		0.34		1.36		1.84
			605		38,934		0.00		30.00		0.36		1.02		1.05
Western			800		1,078		0.00		55.00		1.48		4.70		22.10
			901		153		0.00		250.00		58.17		63.11		3982.66
Silver Zones			902		85		0.23		80.00		21.16		22.61		511.19
			903		136		0.00		70.00		15.77		16.84		283.67
			904		420		0.00		115.00		15.35		19.30		372.68

14-17

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

Table 14-7: Basic Statistics of Composites for Domain 200 (Zone 34)


Data	Count		Minimum		Maximum		Mean		Standard Deviation		Sample Variance
Composite
Ni_ppm		465		0.00		44,365.77		2,081.04		5,701.44		32,506,462.78
Cu_ppm		465		0.00		34,201.35		1,531.24		4,517.75		20,410,052.92
Pt_ppm		465		0.00		8,354.73		185.86		701.62		492,277.56
Pd_ppm		465		0.00		21,913.27		543.45		1,855.97		3,444,638.38
Au_g/t		465		0.00		26.57		0.35		1.64		2.70
Capped Composite
Ni_ppm		465		0.00		35,000.00		2,030.65		5,433.71		29,525,257.63
Cu_ppm		465		0.00		25,000.00		1,452.73		4,172.56		17,410,263.35
Pt_ppm		465		0.00		4,000.00		168.40		563.41		317,430.46
Pd_ppm		465		0.00		8,000.00		481.35		1,399.06		1,957,368.18
Au_g/t		465		0.00		3.00		0.24		0.49		0.24

14-18

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

Table 14-8: Basic Statistics for Calcium Uncapped Composites for All Resource Domains


Zone	Domain	Subdomain		Count		Minimum		Maximum		Mean		Standard Deviation		Sample Variance
	Low		100		73,579		0.00		10.37		1.36		1.16		1.34
			110		3,256		0.01		6.89		0.96		1.01		1.03
			111		3,842		0.00		8.10		0.76		0.79		0.62
			112		3,480		0.01		9.79		1.45		1.01		1.02
	Medium		113		1,832		0.00		8.78		0.83		1.09		1.20
			114		838		0.01		5.70		1.18		0.84		0.71
ODM/17			115		722		0.01		9.07		1.35		1.28		1.65
			mg_110_114		13,970		0.00		9.79		1.04		1.01		1.03
			120		1,935		0.00		6.35		0.77		0.90		0.82
			121		3,044		0.00		8.02		0.62		0.71		0.50
	High		122		3,184		0.01		5.32		1.21		0.94		0.88
			123		448		0.01		4.90		0.79		0.88		0.77
			hg_120_123		8,611		0.00		8.02		0.88		0.89		0.80
Zone 34	Low		200		2473		0.00		25.46		0.29		0.90		0.81
	Low		300		11,648		0.01		8.40		1.04		0.85		0.72
433	Medium		310		2,787		0.01		6.77		0.99		0.69		0.48
	High		320		995		0.00		6.29		0.95		0.64		0.42
HS			400		10,308		0.01		8.79		0.99		0.81		0.66
CAP			500		11,075		0.01		13.29		2.56		2.12		4.50
600 Series			601		53,445		0.01		15.42		1.86		1.54		2.38
			602		30,451		0.00		9.37		1.44		1.08		1.17
			603		52,569		0.00		10.00		1.35		1.06		1.13
			604		83,067		0.00		15.89		1.93		2.00		4.02
			605		35,989		0.00		13.69		1.74		1.91		3.65
Western			800		1,078		0.01		10.03		2.00		1.27		1.62
Silver Zones			901		153		0.01		4.16		1.21		0.90		0.81
			902		115		0.01		3.86		0.77		0.77		0.59
			903		88		0.08		3.79		1.02		0.98		0.97
			904		385		0.01		5.86		0.92		0.98		0.96

14-19

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

Table 14-9: Basic Statistics for Calcium Capped Composites for All Resource Domains


Zone	Domain	Subdomain		Count		Minimum		Maximum		Mean		Standard Deviation		Sample Variance
	Low		100		73,579		0.00		8.50		1.36		1.16		1.34
			110		3,256		0.01		5.50		0.96		1.01		1.01
			111		3,842		0.00		5.00		0.75		0.77		0.59
			112		3,480		0.01		6.00		1.44		0.99		0.99
	Medium		113		1,832		0.00		7.00		0.82		1.07		1.14
			114		838		0.01		3.50		1.18		0.83		0.68
ODM/17			115		722		0.01		5.00		1.34		1.22		1.49
			mg_110_114		13,970		0.00		7.00		1.04		0.99		0.99
			120		1,935		0.00		4.00		0.77		0.89		0.79
			121		3,044		0.00		5.00		0.62		0.68		0.47
	High		122		3,184		0.01		5.00		1.21		0.94		0.88
			123		448		0.01		3.50		0.79		0.85		0.73
			hg_120_123		8,611		0.00		5.00		0.88		0.88		0.78
Zone 34	Low		200		2473		0.01		6.00		0.46		0.95		0.89
	Low		300		11,648		0.01		7.00		1.04		0.84		0.71
433	Medium		310		2,787		0.01		4.00		0.99		0.67		0.44
	High		320		995		0.00		6.29		0.95		0.64		0.42
HS			400		10,308		0.01		8.00		0.99		0.81		0.66
CAP			500		11,075		0.01		9.00		2.56		2.12		4.49
600 Series			601		53,445		0.01		9.00		1.86		1.54		2.37
			602		30,451		0.00		7.00		1.44		1.08		1.16
			603		52,569		0.00		8.00		1.35		1.06		1.13
			604		83,067		0.00		9.50		1.92		2.00		4.01
			605		35,989		0.00		9.00		1.74		1.91		3.63
Western			800		1,078		0.01		7.00		1.99		1.24		1.53
Silver Zones			901		153		0.01		3.00		1.20		0.88		0.77
			902		115		0.01		3.00		0.77		0.74		0.55
			903		88		0.08		3.00		1.01		0.96		0.92
			904		385		0.01		3.00		0.90		0.93		0.86

14-20

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

Table 14-10: Basic Statistics for Sulphur Uncapped Composites for All Resource Domains


Zone	Domain	Subdomain		Count		Minimum		Maximum		Mean		Standard Deviation		Sample Variance
	Low		100		73,579		0.00		14.52		0.91		1.06		1.12
			110		3,256		0.00		11.25		1.24		1.46		2.13
			111		3,842		0.00		10.00		1.22		1.13		1.27
			112		3,480		0.00		10.10		1.08		1.05		1.10
	Medium		113		1,832		0.00		4.84		0.63		0.89		0.80
			114		838		0.01		5.27		1.34		0.91		0.83
ODM/17			115		722		0.00		8.70		1.73		1.32		1.73
			mg_110_114		13,970		0.00		11.25		1.14		1.19		1.42
			120		1,935		0.00		11.76		1.42		1.68		2.83
			121		3,044		0.00		11.54		1.48		1.31		1.71
	High		122		3,184		0.00		10.10		1.43		1.32		1.74
			123		448		0.01		5.28		0.96		1.13		1.29
			hg_120_123		8,611		0.00		11.76		1.42		1.40		1.96
Zone 34	Low		200		2473		0.01		7.48		0.22		0.57		0.32
	Low		300		11,648		0.00		13.37		1.39		1.75		3.05
433	Medium		310		2,787		0.00		11.42		1.54		1.82		3.32
	High		320		995		0.00		13.34		1.66		1.78		3.15
HS			400		10,308		0.00		11.35		1.53		1.41		2.00
CAP			500		11,075		0.01		14.23		1.63		2.13		4.55
600 Series			601		53,445		0.01		14.47		1.41		1.82		3.30
			602		30,451		0.00		13.92		1.00		1.06		1.13
			603		52,569		0.00		13.79		0.94		1.11		1.22
			604		83,067		0.00		13.42		0.49		0.95		0.89
			605		35,989		0.00		11.06		0.47		0.84		0.70
Western			800		1,078		0.01		9.99		1.49		1.37		1.86
Silver Zones			901		153		0.01		3.66		0.88		0.70		0.49
			902		115		0.01		3.60		1.29		0.97		0.94
			903		88		0.01		7.84		2.60		1.92		3.69
			904		385		0.00		3.32		0.67		0.81		0.66

14-21

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

Table 14-11: Basic Statistics for Sulphur Capped Composites for All Resource Domains


Zone	Domain	Subdomain		Count		Minimum		Maximum		Mean		Standard Deviation		Sample Variance
	Low		100		73,579		0.00		10.50		0.91		1.05		1.11
			110		3,256		0.00		9.00		1.23		1.45		2.09
			111		3,842		0.00		6.00		1.21		1.10		1.21
			112		3,480		0.00		5.00		1.07		0.98		0.96
	Medium		113		1,832		0.00		3.50		0.63		0.88		0.78
			114		838		0.01		3.50		1.34		0.91		0.82
ODM/17			115		722		0.00		7.00		1.73		1.31		1.70
			mg_110_114		13,970		0.00		9.00		1.14		1.17		1.36
			120		1,935		0.00		9.00		1.42		1.66		2.75
			121		3,044		0.00		7.00		1.48		1.29		1.65
	High		122		3,184		0.00		8.00		1.43		1.30		1.69
			123		448		0.01		4.00		0.95		1.11		1.22
			hg_120_123		8,611		0.00		9.00		1.42		1.38		1.90
Zone 34	Low		200		2473		0.01		4.00		0.22		0.54		0.29
	Low		300		11,648		0.00		10.00		1.39		1.73		3.01
433	Medium		310		2,787		0.00		10.00		1.54		1.82		3.31
	High		320		995		0.00		9.00		1.64		1.69		2.86
HS			400		10,308		0.00		10.00		1.53		1.41		2.00
CAP			500		11,075		0.01		10.00		1.63		2.12		4.50
600 Series			601		53,445		0.01		12.00		1.41		1.81		3.29
			602		30,451		0.00		9.50		1.00		1.06		1.12
			603		52,569		0.00		10.00		0.94		1.10		1.21
			604		83,067		0.00		10.00		0.49		0.94		0.89
			605		35,989		0.00		8.50		0.47		0.83		0.69
Western			800		1,078		0.01		8.00		1.49		1.34		1.79
Silver Zones			901		153		0.01		2.00		0.84		0.61		0.37
			902		115		0.01		3.00		1.28		0.95		0.89
			903		88		0.01		4.50		2.43		1.66		2.75
			904		385		0.00		3.00		0.67		0.81		0.65

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14.7.2

Variography

Variography was undertaken using GSLib software to characterize the spatial continuity of the metal grade data in each resource domain. Variograms were modelled for gold, silver, calcium, and sulphur for all zones.

The general methodology to calculate and model variograms consists of calculating both directional and isotropic variograms. For each resource domain and for each variable, SRK examined three (3) different spatial metrics: one (1) traditional semivariogram, two (2) traditional correlogram, and three (3) normal score semivariogram.

In general, the correlogram and normal scores transform facilitate the identification of spatial structure, particularly when the traditional variogram shows little continuity. Wherever possible, the traditional variogram was used for modelling; in cases where the traditional variogram was too noisy or unstable, one (1) or a combination of the other three (3) metrics was used to identify the continuity structure.

Generally, the sulphide mineralization at Rainy River strikes at approximately 096º and dips 55º to the south, although local variations do occur. While the majority of the variograms were modelled with two structures, there are a few cases where a third structure was also modelled. In most cases, an exponential model was fitted to the first structure. The remaining structure(s) were often fitted with a spherical model. The down-hole variograms were used to estimate the nugget effect and the ranges in the Z direction (thickness of the domain).

Figure 14-1 presents a sample of the variograms produced by SRK. Variogram parameters are summarized in Table 14-12 to Table 14-15. The variograms for all other resource domains are presented in Appendix E.

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LOGO

Figure 14-5: Examples of Variogram Models for Rainy River Gold Deposit

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Table 14-12: Modeled Gold Variogram Parameters for All Resource Domains Grade Interpolation


Zone	Domain	Subdomain			No. Struct.		Nugget		C1		Mod^‡		RX1		RY1		RZ1		C2		Mod^‡		RX2		RY2		RZ2
	Low		100	*		3		0.2		0.2		2		10		15		10		0.3		2		80		60		70
			110			2		0.2		0.7		2		42		50		8		0.1		1		150		90		50
			111			2		0.2		0.6		2		10		15		5		0.2		1		140		80		25
	Medium		112			2		0.3		0.6		2		15		15		7		0.1		1		100		90		50
ODM/17			113	*		3		0.2		0.5		2		25		0		10		0.1		1		25		90		40
Zone			114			1		0.25		0.75		1		70		70		8
			115			2		0.2		0.65		2		40		40		25		0.15		1		130		130		40
			120			2		0.2		0.6		2		15		15		5		0.2				70		70		25
			121	*		3		0.2		0.65		2		15		15		6		0.05		1		15		60		13
	High		122	*		3		0.2		0.5		2		15		15		4		0.15		1		50		70		30
			123	^†		2		0.2		0.6		2		15		15		5		0.2				70		70		25
Zone 34	Low		200			2		0.15		0.25		1		10		10		10		0.6		1		75		55		35
	Low		300	*		3		0.1		0.4		2		10		30		8		0.25		1		100		45		25
433 Zone	Medium		310			2		0.2		0.6		2		15		35		6		0.2		1		200		60		20
	High		320			2		0.2		0.45		2		10		10		4		0.35		1		60		30		8
HS			400			2		0.2		0.4		2		10		10		5		0.4		1		50		15		20
CAP			500			2		0.2		0.55		2		15		15		10		0.2		1		100		100		70
Western			800			1		0.2		0.8		2		200		200		200
901-904			900			1		0.2		0.8		2		60		60		12

^‡	1 = Exponential; 2 = Spherical

*	For domains of the ODM/17 Zone a third structure was modelled with a spherical function with:

100: C3 = 0.30, RX3 = 500, RY3 =500, and RZ3 = 70

113: C3 = 0.20, RX3 = 110, RY3 =90, and RZ3 = 40

121: C3 = 0.10, RX3 = 140, RY3 =60, and RZ3 = 18

122: C3 = 0.15, RX3 = 160, RY3 =70, and RZ3 = 30

^†	Variogram borrowed from 120

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Table 14-13: Modeled Silver Variogram Parameters for All Resource Domains Grade Interpolation


Zone	Domain	Subdomain			No. Struct.		Nugget		C1		Mod^‡		RX1		RY1		RZ1		C2		Mod^‡		RX2		RY2		RZ2
ODM/17 Zone	Low		100	*		3		0.2		0.4		2		35		30		20		0.15		2		35		30		110
	Medium		110			2		0.2		0.35		2		50		50		20		0.45		1		300		300		40
			111			2		0.2		0.45		2		25		10		15		0.35		1		70		110		70
			112			2		0.2		0.55		2		5		5		5		0.25		1		35		35		35
			113			3		0.2		0.6		2		35		10		20		0.2		1		70		40		40
			114			1		0.2		0.8		2		50		50		20
			115			2		0.2		0.3		2		30		30		30		0.5		1		100		100		40
	High		120			2		0.2		0.4		2		45		150		30		0.4		1		400		300		30
			121			2		0.25		0.45		2		35		60		8		0.3		1		80		100		45
			122			2		0.2		0.5		2		10		10		4		0.3		1		60		60		35
			123			2		0.2		0.35		2		35		35		7		0.45		1		50		50		10
Zone 34	Low		200	*		3		0.2		0.2		2		10		10		15		0.3		2		80		80		40
433 Zone	Low		300	*		3		0.2		0.3		2		5		5		10		0.25		1		60		60		100
	Medium		310	*		3		0.2		0.2		2		45		15		8		0.4		2		45		15		60
	High		320			2		0.2		0.2		2		10		35		20		0.6		1		110		35		20
HS			400	*		3		0.2		0.3		2		10		10		15		0.25		1		10		10		80
CAP			500			2		0.2		0.6		2		25		25		15		0.2		1		300		300		70
Western			800			2		0.2		0.2		1		15		15		15		0.6		1		70		70		70
901-904			900			1		0.2		0.8		2		55		55		12

^‡	1 = Exponential; 2 = Spherical

*	For domains of ODM/17 Zone a third structure was modelled with a spherical function with:

100: C3 = 0.25, RX3 = 300, RY3 = 300, and RZ3 = 110

200: C3 = 0.30, RX3 = 220, RY3 = 220, and RZ3 = 40

300: C3 = 0.30, RX3 = 120, RY3 = 120, and RZ3 = 100

310: C3 = 0.20, RX3 = 25, RY3 = 90, and RZ3 = 60

400: C3 = 0.25, RX3 = 100, RY3 = 75, and RZ3 = 80

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Table 14-14: Modeled Calcium Variogram Parameters for All Resource Domains Grade Interpolation


Zone	Domain	Subdomain			No. Struct.		Nugget		C1		Mod^‡		RX1		RY1		RZ1		C2		Mod^‡		RX2		RY2		RZ2
ODM/17 Zone	Low		100	*		3		0.2		0.4		2		15		15		120		0.4		2		50		50		140
	Medium		110	*		3		0.15		0.35		2		10		15		80		0.2		1		50		15		80
			111	*		3		0.15		0.45		2		10		15		60		0.2		1		70		15		60
			112			2		0.15		0.55		2		15		15		50		0.3		1		200		100		100
			113			2		0.2		0.6		2		25		25		80		0.3		1		130		130		80
			114			1		0.25		0.8		1		70		70		70
			115			2		0.1		0.3		2		130		130		50		0.15		1		130		130		50
	High		120	*		3		0.15		0.4		2		25		25		110		0.25		1		25		120		110
			121	*		3		0.15		0.45		2		10		20		55		0.25		1		80		20		55
			122	*		3		0.15		0.5		2		10		10		40		0.25		1		10		10		100
			123			2		0.2		0.35		2		70		15		5		0.3		1		70		20		15
Zone 34	Low		200			2		0.2		0.2		2		25		25		10		0.3		1		200		200		35
433 Zone	Low		300	*		3		0.1		0.3		2		40		30		30		0.1		1		250		30		30
	Medium		310	*		2		0.2		0.2		2		100		100		80		0.35		1		200		160		80
	High		320			2		0.2		0.2		2		90		90		30		0.35		1		200		200		80
HS			400			2		0.2		0.3		2		40		40		80		0.4		1		60		60		80
CAP			500			2		0.2		0.6		2		35		35		45		0.25		1		180		180		70
Western			800			1		0.2		0.2		2		30		30		15
901-904			900			1		0.2		0.8		2		60		60		25

^‡	1 = Exponential; 2 = Spherical

*	For domains of ODM/17 Zone a third structure was modelled with a spherical function with:

100: C3 = 0.20, RX3 = 500, RY3 = 500, and RZ3 = 140

110: C3 = 0.20, RX3 = 50, RY3 = 120, and RZ3 = 80

111: C3 = 0.20, RX3 = 70, RY3 = 50, and RZ3 = 60

120: C3 = 0.25, RX3 = 120, RY3 = 120, and RZ3 = 110

121: C3 = 0.25, RX3 = 200, RY3 = 160, and RZ3 = 55

122: C3 = 0.25, RX3 = 180, RY3 = 90, and RZ3 = 100

300: C3 = 0.40, RX3 = 250, RY3 = 200, and RZ3 = 200

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Table 14-15: Modeled Sulphur Variogram Considered for All Resource Domains Grade Interpolation


Zone	Domain	Subdomain			No. Struct.		Nugget		C1		Mod^‡		RX1		RY1		RZ1		C2		Mod^‡		RX2		RY2		RZ2
ODM/17 Zone	Low		100			2		0.15		0.75		2		30		30		140		0.1		1		300		300		140
	Medium		110	*		3		0.15		0.45		2		10		25		80		0.2		1		45		15		80
			111			2		0.15		0.45		2		10		10		40		0.4		1		30		30		40
			112			2		0.15		0.6		2		15		15		70		0.25		1		80		80		100
			113	*		3		0.2		0.5		2		70		25		60		0.1		1		130		50		60
			114			1		0.2		0.8		1		70		70		70
			115			1		0.1		0.9		2		100		100		50
	High		120	*		3		0.15		0.35		2		25		25		80		0.25		1		25		90		80
			121			2		0.15		0.6		2		30		20		55		0.25		1		30		40		55
			122	*		3		0.15		0.35		2		10		10		40		0.25		1		10		10		100
			123			1		0.2		0.8		2		110		40		10
Zone 34	Low		200			2		0.2		0.4		2		30		30		10		0.4		1		30		30		35
433 Zone	Low		300			2		0.1		0.4		2		40		25		100		0.5		1		250		170		100
	Medium		310			2		0.2		0.45		2		90		110		70		0.35		1		250		110		70
	High		320			2		0.2		0.45		2		140		90		80		0.35		1		250		140		80
HS			400			2		0.15		0.65		2		40		40		100		0.2		1		100		100		100
CAP			500			2		0.15		0.75		2		10		10		45		0.1		1		80		80		60
Western			800			1		0.2		0.8		2		80		80		15
901-904			900			1		0.2		0.8		2		60		60		25

^‡	1 = Spherical; 2 = Exponential

*	For domains of ODM/17 Zone a third structure was modelled with a spherical function with:

110: C3 = 0.20, RX3 = 45 RY3 = 120, and RZ3 = 80

113: C3 = 0.20, RX3 = 1,300, RY3 = 50, and RZ3 = 200

120: C3 = 0.25, RX3 = 120, RY3 =110, and RZ3 = 80

122: C3 = 0.25, RX3 = 100, RY3 = 100, and RZ3 = 100

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14.8

Block Model and Grade Estimation

14.8.1

Block Model Definition

An unrotated block model was created in Gemcom to cover the entire extent of the Rainy River Gold Project deposit area. Block size was set at 5 m by 5 m by 5 m (Table 14-16).

Criteria used in the selection of block size includes the borehole spacing, composite assay length, consideration for the potential size of the smallest mining unit and the geometry of the modelled auriferous Zones.

Table 14-16: Rainy River Gold Project Block Model Parameters


Direction	Origin	Size (m)		Minimum*		Maximum*		Number of Cells
East-West			5		423,700		427,075		675
North-South			5		5,408,750		5,411,125		475
Vertical			5		(-)1,200		450		330

*	UTM NAD83 datum, Zone 15

14.8.2

Grade and Specific Gravity Estimation

Metal grades were estimated using ordinary kriging as the principal estimator separately in each domain from capped composite data within that domain. Grades in Domains 601 to 605 were estimated using an inverse distance algorithm similarly to that in the previous mineral resource model. In a sensitivity study undertaken by SRK in 2010, it was determined that estimation for the Rainy River Gold Project domains is insensitive to slight variations of estimation parameters. Grade interpolation was completed in two (2) or three (3) successive passes considering the estimation parameters summarized in Table 14-17 and the search neighbourhoods summarized in Table 14-18.

The search neighbourhoods used for calcium and sulphur interpolation are tabulated in Table 14-19 and Table 14-20, whereas that for Domain 200 (Zone 34) is summarized in Table 14-21. The first estimation pass considered search neighborhood adjusted to 95% of the variogram sill. The size of the search ellipse was doubled for the second estimation pass. A third

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estimation pass was used for Domains 114, 300, and 400 with a search distance set at three (3) times the first pass range in order to populate all the blocks in the entire domain.

Grade interpolation within the 600 domains has been further restricted by limiting the influence of high grades to half the distance of the first pass ellipsoidal search. Considering the higher uncertainty in gold mineralization continuity within the 600 series, a grade restriction of 8 g/t gold was applied.

For the ODM/17, 433, HS and CAP domains a specific gravity value was estimated in each model block using an inverse distance algorithm in a single pass with the following criteria:

•

A minimum of 4 and a maximum of 20 composites;

•

A maximum of 3 composites per borehole; and

•

A spherical search radius of 500 m.

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Table 14-17: Resource Estimation Parameters


Interpolation Parameters	1^stPass (Indicated)		2^ndPass (Inferred)		3^rdPass (Inferred)
Domains 100-500, 800 and 900
Interpolation Method		Ordinary Kriging		Ordinary Kriging		Ordinary Kriging
Search Type		Octant		Ellipsoidal		Ellipsoidal
Minimum Number of Octants		2		—		—
Maximum Number of Composites per Octant		5		—		—
Minimum Number of Composites		3		2		2
Maximum Number of Composites		8		12		12
Maximum Number of Composite per Borehole		2		—		—
Domains 601 to 605
Interpolation Method		Inverse Distance Power 2		Inverse Distance Power 2		—
Search Type		Ellipsoidal		Ellipsoidal		—
Minimum Number of Composites		3		2		—
Maximum Number of Composites		10		15		—
Maximum Number of Composite per Hole		2		—		—
Domain 200 (Pt, Pd, Ni and Cu)
Interpolation Method		Ordinary Kriging		Ordinary Kriging		—
Search Type		Octant		Ellipsoidal		—
Minimum Number of Octants		2		—		—
Maximum Number of Composites per Octant		5		—		—
Minimum Number of Composites		3		2		—
Maximum Number of Composites		10		12		—
Maximum Number of Composite per Hole		2		—		—

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Table 14-18: Search Neighbourhoods Used for Gold and Silver Estimation


	Domain		Rotation*						1^stPass						2^ndPass						3^rdPass
			Rotation*						Search Ranges						Search Ranges						Search Ranges
			Az		Dip		Plunge		X		Y		Z		X		Y		Z		X		Y		Z
Au/Ag		100		-110		-40		132		200		100		50		200		200		100
		110		-120		-40		122		100		60		35		200		120		70
		111		-105		-40		145		95		55		20		190		110		40
		112		-110		-40		132		70		60		35		140		120		70
		113		-120		-40		122		75		60		30		150		120		60
		114		-130		-40		112		50		50		10		100		100		20		150		150		30
		115		-120		-40		122		90		90		30		180		180		60
		120		-120		-40		122		55		55		25		110		110		50
		121		-110		-40		132		95		40		15		190		80		30
		122		-115		-40		127		110		50		25		220		100		50
		123		-120		-40		55		55		25		110		110		50		55
		200		45		56		-55		75		55		35		150		110		70
		300		20		50		-70		70		40		20		140		80		40		175		100		50
		310		25		50		-65		135		40		15		270		80		30
		320		20		45		-75		60		30		20		120		60		20
		400		10		50		-80		50		20		20		100		40		40		150		60		60
		500		15		55		-75		70		70		50		140		140		100
		601		78		-47		55		60		60		32		120		120		64
		602		78		-47		55		120		120		45		240		240		90
		603		78		-47		55		200		200		20		200		200		20
		604		78		-47		55		200		200		20		200		200		20
		605		78		-47		55		170		110		60		340		220		120
		800		0		0		0		130		130		130		260		260		260
		901		40		55		-50		60		60		12		120		120		24
		902		20		45		-70		60		60		12		120		120		24
		903		5		55		-85		60		60		12		120		120		24
		904		20		60		-70		60		60		12		120		120		24

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Table 14-19: Search Neighbourhoods Used for Calcium


	Domain		Rotation*						1^stPass						2^ndPass						3^rdPass
			Rotation*						Search Ranges						Search Ranges						Search Ranges
			Az		Dip		Plunge		X		Y		Z		X		Y		Z		X		Y		Z
Ca		100		-110		-40		132		200		200		50		200		200		100
		110		-120		-40		122		50		120		80		100		240		180
		111		-105		-40		145		70		50		60		140		100		120
		112		-110		-40		132		130		60		75		260		120		150
		113		-120		-40		122		130		130		80		260		260		160
		114		-130		-40		112		70		70		70		140		140		140
		115		-120		-40		122		130		130		50		260		260		100
		120		-120		-40		122		90		80		90		180		160		180
		121		-110		-40		132		100		120		45		200		240		90
		122		-115		-40		127		110		50		70		220		100		140
		123		-120		-40		122		70		20		15		140		40		30		210		60		45
		200		45		56		-55		130		130		35		260		260		70
		300		20		50		-70		180		140		140		360		280		280
		310		25		50		-65		140		110		65		280		220		130
		320		20		45		-75		140		140		55		280		280		110
		400		10		50		-80		60		60		80		120		120		160
		500		15		55		-75		110		110		70		220		220		140
		601		78		-47		55		60		60		32		120		120		64
		602		78		-47		55		120		120		45		240		240		90
		603		78		-47		55		200		200		20		200		200		20
		604		78		-47		55		200		200		20		200		200		20
		605		78		-47		55		170		110		60		340		220		120
		800		0		0		0		30		30		15		60		60		30		120		120		60
		901		40		55		-50		60		60		25		120		120		50
		902		20		45		-70		60		60		25		120		120		50
		903		5		55		-85		60		60		25		120		120		50
		904		20		60		-70		60		60		25		120		120		50

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Table 14-20: Search Neighbourhoods Used for Sulphur


	Domain		Rotation*						1^stPass						2^ndPass						3^rdPass
			Rotation*						Search Ranges						Search Ranges						Search Ranges
			Az		Dip		Plunge		X		Y		Z		X		Y		Z		X		Y		Z
S		100		-110		-40		132		100		100		50		200		200		200
		110		-120		-40		122		45		120		80		90		240		180
		111		-105		-40		145		30		30		40		60		60		80		90		90		120
		112		-110		-40		132		80		80		100		160		160		200
		113		-120		-40		122		130		50		100		260		100		200
		114		-130		-40		112		70		70		70		140		140		140
		115		-120		-40		122		100		100		50		200		200		100
		120		-120		-40		122		80		80		75		160		160		150
		121		-110		-40		132		30		40		55		60		40		110
		122		-115		-40		127		100		100		100		200		200		200
		123		-120		-40		122		110		40		10		220		80		20
		200		45		56		-55		30		30		35		60		60		70
		300		20		50		-70		170		120		75		340		240		150
		310		25		50		-65		160		110		70		320		220		140
		320		20		45		-75		170		100		65		320		200		130
		400		10		50		-80		100		100		100		200		200		200
		500		15		55		-75		80		80		60		160		160		120
		601		78		-47		55		60		60		32		120		120		64
		602		78		-47		55		120		120		45		240		240		90
		603		78		-47		55		200		200		20		200		200		20
		604		78		-47		55		200		200		20		200		200		20
		605		78		-47		55		170		110		60		340		220		120
		800		0		0		0		80		80		15		160		160		30		160		160		60
		901		40		55		-50		60		60		25		120		120		50
		902		20		45		-70		60		60		25		120		120		50
		903		5		55		-85		60		60		25		120		120		50
		904		20		60		-70		60		60		25		120		120		50

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Table 14-21: Search Parameters Used for Grade Estimation in Zone 34


	Domain		1^stPass Search Ranges						2^ndPass Search Ranges
	Domain		X		Y		Z		X		Y		Z
Pt		200		75		75		18		150		150		35
Pd				75		75		18		150		150		35
Ni				35		35		23		70		70		45
Cu				38		38		18		75		75		35

14.9

Model Validation and Sensitivity

The mineral resource model was validated by visually comparing block estimates to informing borehole data on section by section and elevation by elevation basis (as in Figure 14-6). Four (4) representative cross-sections across the Rainy River Gold Project showing block model gold grades in relation to geology zones, conceptual pit outline and informing borehole data are presented in Appendix F.

Quantile-quantile plots comparing block and declustered capped composite data were also constructed for each Zone (Appendix G). These plots also show the usual smoothing effect of kriging, particularly at higher grades. SRK is satisfied that the current block model that fairly represents the validated Rainy River drill data considered for the resource estimation.

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LOGO

Figure 14-6: Cross-Section 425,775E Comparing Blocks Populated

with Gold-Grades and Informing Data

SRK has also carried out the comparative analyses between ordinary kriging estimates and those estimated using an inverse distance (power of 2) estimator. Both estimation techniques show comparative results.

14.10

Mineral Resource Classification

Mineral resources were classified according to the CIM Definition Standards for Mineral Resources and Mineral Reserves (November 2010) by Dorota El-Rassi, P.Eng. (PEO #100012348), and Glen Cole, P.Geo. (APGO #1416), appropriate independent qualified persons for the purpose of National Instrument 43-101.

The mineral resources are classified primarily based on the basis of a block’s distance from the nearest informing composites and on variography results. Classification is based on gold data

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alone. Generally, an Indicated classification is assigned to blocks estimated during the first estimation pass using 95% of the variogram sill; whereas an Inferred classification is assigned to all other blocks estimated during the second or third estimation pass. SRK also considered the confidence in the geological interpretation during the classification process.

As a result of the infill drilling program completed in 2012 and the increased confidence in the geological and grade continuity, SRK considers that a Measured classification can be assigned to certain parts of the OMD/17 Zone (domains 100 and 300) where the drilling information and density is sufficient to confirm the geological and grade continuity within the meaning of the CIM Definition Standards for Mineral Resources and Mineral Reserves. The parameters used to categorize the Measured blocks are summarized in Table 14-22.

The classification strategy also considered the geological setting as well as what impact additional drill data would have on the shape of the modelled resource domain and the confidence in the grade continuity. All resource blocks within domains 601 to 605 are estimated at a low level of confidence and therefore have been assigned an Inferred classification.

Table 14-22: Search Parameters Used to Code the Measured Blocks


Interpolation Parameters		Measured
Domains 100 and 300
Interpolation Method		Ordinary Kriging
Search Type		Octant (25x25x25)
Minimum Number of Octants		3
Maximum Number of Composites per Octant		4
Minimum Number of Composites		5
Maximum Number of Composites		8
Maximum Number of Composite per Borehole		2

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14.11

Mineral Resource Statement

CIM Definition Standards for Mineral Resources and Mineral Reserves (November 2010) defines a mineral resource as:

“A concentration or occurrence of diamonds, natural solid inorganic material, or natural solid fossilized organic material including base and precious metals, coal, and industrial minerals in or on the Earth’s crust in such form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction. The location, quantity, grade, geological characteristics and continuity of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge.”

The “reasonable prospects for economic extraction” requirement generally implies that the quantity and grade estimates meet certain economic thresholds and that the mineral resources are reported at an appropriate cut-off grade that takes into account extraction scenarios and processing recoveries. SRK considers that portions of the Rainy River gold deposits are amenable for open pit extraction, while other parts of the deposits could be extracted using an underground mining method.

To assist with determining which portions of the modelled mineralization show “reasonable prospect for economic extraction” from an open pit, and to assist with selecting reasonable reporting assumptions, SRK used a pit optimizer to develop conceptual open pit shells using the assumptions summarized in Table 14-23. The block model quantities and grade estimates were also reviewed to determine the portions of the modelled mineralization having “reasonable prospects for economic extraction” from an underground mine, based on parameters summarized in Table 14-24.

The reader is cautioned that the results from the pit optimization are used solely for the purpose of assessing those portions of the block models that show “reasonable prospects for economic extraction” by an open pit and do not represent and attempt to evaluate mineral reserves. Mineral reserves can only be estimated based on the results of an economic evaluation as part of a

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preliminary feasibility study or a feasibility study. As such, no mineral reserves have been estimated by SRK. There is no certainty that all or any part of the mineral resource will be converted into mineral reserve.

Table 14-23: Conceptual Assumptions Considered for Open Pit Resource Reporting


Parameter		Assumption
Pit Wall Angle		Average 48°
Mining Cost (Ore and Waste)		CAD $2.00/t rock
Process Cost Including G & A Costs		USD $7.25/t
Process Recovery		88% gold, 75% silver
Assumed Process Rate		32,000 tpd from open pit and underground
Metal Price		USD $1,100/oz. gold and USD $22.50/oz. silver and USD $1,600/oz. gold and USD $22.50/oz. silver
Mining Dilution and Losses		10.0%

Table 14-24: Conceptual Assumptions Considered for Underground Resource Reporting


Parameter		Assumption
Mining Costs		CAD $55.00/t rock
Assumed Process Rate		32,000 tpd from open pit and underground
Assumed Mining Rate		2,500 tpd
Process Cost Including G & A Costs		USD $7.25/t
Process Recovery		90% gold, 75% silver
Metal Price		USD $1,100/oz. gold and USD $22.50/oz. silver
Mining Dilution		20.0%
Mining Recoveries		100%

SRK considers that material above an elevation of -150 metres above sea level (“masl”) offers reasonable prospects for economic extraction from an open pit.

In order to prepare the Mineral Resource Statement for the Rainy River Gold Project, SRK considered two (2) conceptual pit shells defined using two (2) gold prices USD $1,100 and

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USD $1,600 per ounce. After review of optimization results, SRK subdivided the block model in four (4) areas for reporting mineral resources (shown in Figure 14-7):

1.	Open pit, Measured, Indicated, and Inferred mineral resources reported inside the USD $1,100 conceptual pit shell above an elevation of -150 masl;

2.	Open pit, Indicated and Inferred mineral resources reported outside the USD $1,100 conceptual shell but inside the USD $1,600 conceptual pit shell above the elevation of -150 masl;

3.	Open pit, Inferred mineral resources reported outside the USD $1,600 conceptual pit shell above the elevation of -150 masl; and

4.	Underground, Measured, Indicated and Inferred mineral resources are reported below an elevation of -150 masl.

SRK is satisfied that the material outside the USD $1,600 conceptual pit shells, but above the final depth of this shell (-150 masl elevation), meets the minimum requirements for a mineral resource amenable for open pit extraction; however, considering the uncertainty, all blocks above the elevation of -150 masl and outside the USD $1,600 conceptual pit envelopes are reclassified as Inferred mineral resources.

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LOGO

Figure 14-7: Schematic Vertical Section Illustrating Criteria Considered for Preparing the Mineral Resource Statement for the Rainy River Gold Project (View Looking East)

The mineral resources may be affected by subsequent assessments of mining, environmental, processing, permitting, taxation, socio-economic, and other factors. There is insufficient information at this early stage of the study to assess the extent to which the resources will be affected by these factors.

The Rainy River Gold Project contains gold-rich polymetallic sulphide mineralization of hydrothermal origin, which has been overprinted by magmatic copper-nickel sulphide mineralization enriched in platinum group metals (Zone 34). Copper and nickel grades were estimated in the block model, however, these metals do not contribute significantly to the overall value of the gold-rich sulphide mineralization. Hence, the Mineral Resource Statement for the Rainy River Gold Project is reported on the basis of gold and silver grades only. The mineral resources for Zone 34 and the Silver Zone are reported separately since they contain different metal characteristics.

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The Mineral Resource Statement is reported at two (2) gold cut-off grades considering the most likely extraction scenario. Open pit mineral resources are reported at a cut-off grade of 0.35 g/t gold, whereas underground mineral resources are reported at a cut-off grade of 2.50 g/t gold. Cut-off grades are based on a long term gold price of USD $1,100 per ounce and a gold, metallurgical recovery of 88% and 90% for open pit and underground mineral resources, without considering revenues from other metals. The consolidated Mineral Resource Statement for the Rainy River Gold Project gold zones is summarized in Table 14-25 and excludes gold mineralization from Zone 34 and the Silver Zone. The effective date of this eighth Mineral Resource Statement prepared for the Rainy River Gold Project is October 10, 2012.

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Table 14-25: Consolidated Mineral Resource Statement*, Rainy River Gold Project, Ontario, SRK Consulting (Canada) Inc., October 10, 2012


Category	Quantity ‘000 t		Grade				Metal
Category	Quantity ‘000 t		Au g/t		Ag g/t		Au ‘000 oz.		Ag ‘000 oz.
In Pit Mineral Resources**
Measured		27,550		1.32		1.90		1,168		1,681
Indicated		112,271		1.11		2.51		4,012		9,048
Measured and Indicated		139,821		1.15		2.39		5,180		10,728
Inferred		19,353		0.88		1.40		550		870
Out of Pit Mineral Resources**
Indicated		14,466		0.80		3.84		373		1,785
Inferred		73,555		0.68		2.53		1,610		5,980
Underground Mineral Resources**
Measured		88		4.97		2.76		14		8
Indicated		4,148		4.50		6.12		600		816
Measured and Indicated		4,236		4.50		6.05		614		824
Inferred		897		4.18		4.63		120		134
Combined Mineral Resources: In Pit, Out of Pit and Underground**
Measured		27,638		1.33		1.90		1,182		1,689
Indicated		130,885		1.18		2.46		4,985		11,649
Measured and Indicated		158,523		1.21		2.62		6,167		13,338
Inferred		93,805		0.75		2.32		2,280		6,984

Mineral resources are reported by relative conceptual pit shells. On average, the open pit extends to an elevation of 500 m below surface. Mineral resources are not mineral reserves and do not have a demonstrated economic viability. All figures are rounded to reflect the relative accuracy of the estimate. Figures may not add due to rounding. All assays have been capped where appropriate. Qualified Persons - The mineral resource statement was prepared by Dorota El-Rassi, P.Eng. (APEO #100012348) and Glen Cole, P.Geo. (APGO #1416), of SRK, both “independent qualified persons” as that term is defined in National Instrument 43-101. Rainy River’s exploration program in Richardson Township is being supervised by Kerry Sparkes, P.Geo. (APEGBC #25261), Vice-President, Exploration and a Qualified Person as defined by National Instrument 43- 101. The Company continues to implement a rigorous QA/QC program to ensure best practices in sampling and analysis of drill core. Mineral resource estimates may be materially affected by environmental, permitting, legal, title, taxation, sociopolitical, marketing, and other relevant issues.

Open pit mineral resources are reported at a cut-off grade of 0.35 g/t gold, underground mineral resources are reported at a cut-off grade of 2.5 g/t gold based on a gold price of USD $1,100 per ounce, a silver price of USD $22.50 per ounce, a foreign exchange rate of CAD $1.10 to USD $1.00, gold recovery of 88% for open pit resources and 90% for underground resources with silver recovery at 75%.

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Mineral resources for Zone 34 and the Silver Zone are reported in Table 14-26 and Table 14-27, respectively.

Table 14-26: Mineral Resources* for the 34 Zone (Domain 200), Rainy River Gold Project, Ontario,

SRK Consulting (Canada) Inc., October 10, 2012


			Grade										Metal
Category	Quantity ‘000 t		Au g/t		Pt g/t		Pd g/t		Ni ppm		Cu ppm		Au ‘000 oz.		Pt ‘000 oz.		Pd ‘000 oz.		Ni t		Cu t
Open Pit Mineral Resources**
Indicated		145		0.66		0.26		0.67		3,477		2,414		3.10		1.21		3.12		503		350

Excluded from the main Mineral Resource Statement. Mineral resources are reported in relation to conceptual pit shells. Mineral resources are not mineral reserves and do not have a demonstrated economic viability. All figures are rounded to reflect the relative accuracy of the estimate. All composites have been capped where appropriate.

**	Open pit mineral resources are reported at a cut-off grade of 0.35 g/t gold. Cut-off grades are based on a price of USD $1,100 per ounce of gold, exchange rate of CAD $1.10 to USD $1.00, and gold recovery of 88%.

Table 14-27: Mineral Resources* for the Silver Zone (Domain 901), Rainy River Gold Project,

SRK Consulting (Canada) Inc., October 10, 2012

Grade

Metal

Category

Quantity
‘000 t

Au
g/t

Ag
g/t

AuEq
g/t

Au
‘000 oz.

Ag
‘000 oz.

AuEq
‘000 oz.

Open Pit Mineral Resources**

Indicated

1,534

0.41

25.03

0.85

20.00

1,235

41.95

Open pit mineral resources are reported at a cut-off grade of 0.35 g/t gold equivalent. Gold equivalent grade is based on a gold price of USD $1,100 per ounce, a silver price of USD $22.50 per ounce, a foreign exchange rate of CAD $1.10 to USD $1.00, gold recovery of 88%, and silver recovery of 75%.

Mineral resources reported by zone and classification are summarized in Table 14-28 and Table 14-29 for open pit and underground areas, respectively.

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Table 14-28: Open Pit Mineral Resources*, Rainy River Gold Project, Ontario,

SRK Consulting (Canada) Inc., October 10, 2012


	Measured						Indicated						Inferred
Domain	Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.		Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.		Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.
Within Pit Shells
ODM/17		23,004		1.32		979		74,914		1.24		2,984
433		4,546		1.30		189		10,944		1.21		425		153		0.88		4
HS								10,234		0.79		260		4,677		0.70		105
CAP								16,180		0.66		343
Western														2,720		1.39		121
601														1,826		1.33		78
602														3,505		0.77		87
603														3,035		0.78		77
604														2,830		0.67		61
605														606		0.81		16

Total		27,550		1.32		1,168		112,271		1.11		4,012		19,353		0.88		550

Outside Pit Shells
ODM/17								8,382		0.86		231
433								1,610		1.00		52		18		0.51		0
HS								565		0.70		13		1,381		0.65		29
CAP								3,908		0.62		78
Western														1,651		0.65		35
601														1,420		0.77		35
602														863		0.65		18
603														1,785		0.72		42
604														1,426		0.73		33
605														362		0.73		8

Total								14,466		0.80		373		8,905		0.70		200

*	Reported at a cut-off grade of 0.35 g/t gold based on a gold price USD $1,100 per ounce gold and assuming a gold metallurgical recovery of 85%. Other metals not considered.

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Table 14-29: Underground Mineral Resources*, Rainy River Gold Project, Ontario,

SRK Consulting (Canada) Inc., October 10, 2012


Domain	Measured						Indicated						Inferred
Domain	Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.		Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.		Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.
ODM/17		88		4.97		14		3,841		4.53		560		3		2.67		0
433								285		3.93		36		58		4.56		8
HS								22		5.48		4		105		3.49		12
CAP														80		3.16		8
Western
601														98		7.33		23
602														30		3.39		3
603														310		4.06		40
604														207		3.68		24
605														7		2.96		1

Total		88		4.97		14		4,148		4.50		600		897		4.18		120

*	Reported at a cut-off grade of 2.50 g/t gold based on a gold price USD $1,100 per ounce gold and assuming a gold metallurgical recovery of 90%. Other metals not considered.

14.12

Grade Sensitivity Analysis

The mineral resources of the Rainy River Gold Project are highly sensitive to the selection of a reporting cut-off grade. To illustrate this sensitivity, grade tonnage curves are presented in Figure 14-8, block model quantities and grade estimates are presented in Table 14-30 at various cut-off grades. Table 14-31 and Table 14-32 show the sensitivity of potential open pit and underground material to the gold cut-off grade. The reader is cautioned that the figures in these tables should not be misconstrued as a Mineral Resource Statement. The figures are only presented to show the sensitivity of the block model estimates to the selection of a reporting cut-off grade.

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LOGO

Figure 14-8: Rainy River Gold Project Global Grade Tonnage Curves

(Open Pit and Underground Material Combined)

Table 14-30: Global Block Model Quantities and Grade Estimates* at Various Cut-Off Grades


Cut-off	Measured						Indicated						Inferred
Au g/t	Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.		Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.		Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.
0.1		50,158		0.85		1,362,973		520,100		0.53		2.01		10,569,893		0.98		45,369,729
0.2		38,936		1.05		1,309,985		361,173		0.70		2.33		6,347,004		1.50		22,478,552
0.3		31,127		1.25		1,248,211		247,068		0.91		2.64		4,686,026		1.82		14,796,673
0.35		28,275		1.34		1,218,545		207,404		1.02		2.79		4,172,207		1.93		12,565,736
0.4		25,861		1.43		1,189,547		176,956		1.13		2.93		3,573,155		2.14		10,528,605
0.5		21,962		1.60		1,133,412		134,872		1.35		3.17		2,586,912		2.57		6,958,957
0.6		18,800		1.78		1,077,744		106,571		1.56		3.39		2,088,761		2.95		5,296,814
0.7		16,188		1.97		1,023,316		86,726		1.77		3.58		1,740,379		3.30		4,145,238
0.8		14,067		2.15		972,315		72,355		1.97		3.75		1,462,223		3.55		3,146,189
0.9		12,313		2.34		924,491		61,605		2.17		3.90		1,310,952		3.82		2,700,457
1.0		10,818		2.53		878,921		53,024		2.37		4.04		1,197,254		4.08		2,394,168
1.2		8,419		2.94		794,531		40,770		2.75		4.27		1,038,254		4.49		1,977,952

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Cut-off	Measured						Indicated						Inferred
Au g/t	Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.		Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.		Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.
1.4		6,687		3.36		722,463		32,340		3.13		4.43		926,838		4.78		1,692,176
1.6		5,431		3.79		662,145		26,186		3.52		4.54		819,326		5.09		1,438,372
1.8		4,485		4.23		610,585		21,610		3.91		4.60		702,696		5.63		1,198,527
2.0		3,784		4.67		567,902		18,080		4.30		4.62		639,341		5.78		1,038,111
2.2		3,236		5.10		530,955		15,399		4.68		4.73		588,751		5.97		927,067
2.4		2,845		5.49		502,009		13,257		5.07		4.84		536,157		6.06		802,176
2.5		2,676		5.68		488,739		12,382		5.25		4.90		512,845		6.04		742,265
2.6		2,516		5.88		475,612		11,589		5.44		4.96		489,712		6.01		684,836
2.8		2,243		6.27		451,997		10,249		5.80		5.08		452,402		6.09		609,788
3.0		2,028		6.62		431,896		9,144		6.15		5.17		425,955		6.11		555,346
3.5		1,623		7.47		389,678		7,084		6.99		5.43		311,769		7.88		428,205
4.0		1,347		8.23		356,547		5,645		7.82		5.69		264,577		8.16		340,179
5.0		976		9.67		303,383		3,879		9.36		5.93		190,411		7.79		195,163
10.0		305		15.81		155,292		1,057		16.29		5.90		50,313		6.47		24,168

*	The reader is cautioned that the figures in this table should not be misconstrued with a Mineral Resource Statement. The figures are only presented to show the sensitivity of the block model estimates to the selection of a cut-off grade.

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Table 14-31: Block Model Quantities and Grade Estimates* at Selective

Cut-off Grades Potential Open Pit Mining Material


Category	Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.		Grade Ag g/t		Ag Metal ‘000 oz.
Cut-off Grade 0.30 g/t Au
Measured		30,333		1.23		1,197		1.84		1,795
Indicated (in pit)		131,702		1.00		4,214		2.40		10,150
Indicated (ex. pit)		18,276		0.70		413		3.56		2,093
Measured & Indicated		180,311		1.01		5,824		2.42		14,038
Inferred (in-pit)		23,799		0.78		596		1.32		1,007
Inferred (ex-pit)		97,406		0.59		1,857		2.28		7,154
Cut-off Grade 0.40 g/t Au
Measured		25,189		1.41		1,140		1.95		1,579
Indicated (in pit)		97,431		1.22		3,834		2.61		8,161
Indicated (ex. pit)		11,628		0.91		339		4.10		1,534
Measured & Indicated		134,248		1.23		5,313		2.62		11,274
Inferred (in-pit)		16,265		0.98		513		1.48		774
Inferred (ex-pit)		55,103		0.78		1,387		2.87		5,081

*	The reader is cautioned that the data presented in this table should not be misconstrued as a Mineral Resource Statement. The figures are only shown to illustrate the sensitivities of the block model quantities and grade estimates to the selection of a cut-off grade.

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Table 14-32: Block Model Quantities and Grade Estimates* at Selected

Cut-off Grades – Potential Underground Mining Material


Category	Quantity ‘000 t		Grade Au g/t		Au Metal ‘000 oz.		Grade Ag g/t		Ag Metal ‘000 oz.
Cut-off Grade 2.0 g/t Au
Measured		113		4.37		16		2.67		10
Indicated		6,773		3.67		800		5.37		1,169
Measured & Indicated		6,886		3.69		816		5.33		1,179
Inferred		1,417		3.47		158		4.13		188
Cut-off grade 3.0 g/t Au
Measured		69		5.58		12		2.96		7
Indicated		3,067		5.21		513		6.82		672
Measured & Indicated		3,136		5.22		526		6.73		679
Inferred		639		4.77		98		4.97		102

*	The reader is cautioned that the data presented in this table should not be misconstrued as a Mineral Resource Statement. The figures are only shown to illustrate the sensitivities of the block model quantities and grade estimates to the selection of a cut-off grade.

Figure 14-9 presents a plan showing the distribution of the open pit mineral resources relative to the USD $1,100 conceptual pit outline.

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LOGO

Figure 14-9: Distribution of Open Pit Mineral Resources Relative to the Conceptual Pit Outline

14.13

Previous Mineral Resource Estimates

A comparison between the February 24, 2012 and the October 10, 2012 Mineral Resource Statements is shown in Table 14-33. The additional 237 boreholes drilled in 2012 have resulted in an overall increase in the contained gold metal. It should be noted that a significant portion of this additional mineral resource is classified as Indicated. The proportion of Measured and Indicated has increased significantly at the expense of Inferred mineral resources.

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Table 14-33: Comparison of February 2012 and October 2012 Mineral Resource Statements


	Quantity			Grade (g/t)						Contained Metal (oz.)
Classification	(tonnes)			Gold			Silver			Gold			Silver
Open Pit
Measured		19	%		2	%		-5	%		5	%		7	%
Indicated		2	%		3	%		13	%		3	%		9	%
Measured & Indicated		5	%		3	%		10	%		6	%		10	%
Inferred		7	%		-3	%		13	%		5	%		6	%
Underground
Measured		-1	%		7	%		8	%		6	%		7	%
Indicated		40	%		4	%		11	%		24	%		37	%
Measured & Indicated		39	%		4	%		11	%		19	%		37	%
Inferred		-27	%		1	%		-34	%		-28	%		-57	%

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15.

MINERAL RESERVE ESTIMATE

As defined by the Canadian Institute of Mining, Metallurgy and Petroleum within the CIM Definition Standards on Mineral Resources and Mineral Reserves (CIM Special Volume 56), the definition of a mineral reserve is as follows:

“A Mineral Reserve is the economically mineable part of a Measured or Indicated Mineral Resource demonstrated by at least a Preliminary Feasibility Study. This Study must include adequate information on mining, processing, metallurgical, economic and other relevant factors that demonstrate, at the time of reporting, that economic extraction can be justified. A Mineral Reserve includes diluting materials and allowances for losses that may occur when the material is mined.”

For studies at the Prefeasibility and Feasibility levels, the CIM guidelines require that only material categorized as Measured or Indicated Resources be classified as a reserve.

15.1

Open Pit Mining

BBA was responsible for the design of the open pit mine and the evaluation of the associated capital and operating costs. Surface mining of the Rainy River gold deposit will follow the standard practice of an open pit operation, with a conventional drill and blast, load and haul cycle using a drill/truck/shovel mining fleet. The overburden and waste rock material will be hauled to the overburden and waste disposal areas near the pit. The run-of-mine ore will be drilled, blasted and loaded by hydraulic shovels and delivered by large mining trucks to the primary crusher or stockpiles.

15.1.1

Resource Block Model

The mining engineering work required for the Study, such as the pit optimization, engineered pit design, mine planning and economic analysis, is based on the resource Block Model prepared by SRK and delivered to BBA on October 1, 2012. The model was transferred by BBA into the mining software, MineSight, for the open pit portion of the Study. The unit block size in the model is

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5 m x 5 m x 5 m. Rainy River provided BBA with digitized topographical and bedrock mapping data. The UTM NAD83 coordinate system was used.

The following data was provided by SRK in the model:

•

UTM coordinates;

•

Rock type (ODM/17, Zone 433, HS Zone, CAP Zone, etc.);

•

Density;

•

Au (gold in g/t);

•

Ag (silver in g/t);

•

Percent (% of the block within a modeled wireframe);

•

OP/UG (Open pit or underground classification, not used);

•

Cat (Categories: Measured, Indicated or Inferred);

•

Pt (platinum, not used);

•

Pd (palladium, not used);

•

Ni (nickel, not used);

•

Cu (copper, not used);

•

S (sulphur, not used);

•

Ca (calcium, not used);

•

NP (acid generation classification – acid neutralizing capacity); and

•

NPR (acid generation classification – net potential ratio).

Additional variables were added to BBA’s MineSight model in order to perform required calculations such as equivalent gold, and to determine block value.

Following the import of the Block Model into MineSight, a verification of the total mineral resources by category was performed in order to insure conformity with the results provided by SRK.

15.1.1.1

Model Surfaces

In addition to the block model file, two (2) surface files were provided to BBA. Both files were provided in the same UTM coordinate system as the block model:

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•

Topography surface; and

•

Overburden surface (interface bedrock/overburden).

The two (2) files provide information about the portions of the block model that are actually above surface, or are below either the topographic surface, in the overburden region, or in bedrock.

The overburden surface that was provided is important for understanding the large variability of overburden thickness in the pit area. The overburden thickness reaches a maximum of approximately 50 m in the centre region of the pit. Also indicated in Figure 15-1 is the final pit shell outline, which will be developed later in this section.

LOGO

Figure 15-1: Isopach Mapping of Overburden Thickness

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15.1.1.2

Specific Gravity

The specific gravity for the mineralized blocks as coded in the SRK Block Model, range from 2.80 t/m³ to 2.94 t/m³ and average 2.85 t/m³. The recommended in-situ overburden density (i.e., material that is at least 50% above the bedrock surface and below the topographic surface) is 1.80 t/m³. Blocks in the model that were not coded as ore or overburden were defaulted as waste. These blocks have a density coding of 2.80 t/m³.

15.1.1.3

Model Recoveries

The model gold and silver recoveries were determined from metallurgical testwork results presented in Section 13 of this report. The silver recovery was determined to a constant value of 64.1%.

An example of Au recovery values is shown in Table 15-1, based on Au grade and ore types in the model.

Table 15-1: Au Recovery by Grade and Ore Type


Head Grade (g/t Au)	Au Recovery in Non-CAP Zones (%)		Au Recovery in CAP Zone (%)
0.2		77.3		74.1
0.4		83.9		74.5
0.6		86.8		74.8
0.8		88.5		74.9
1		89.7		75.1
1.2		90.6		75.2
1.4		91.3		75.3

It is important to note that the recovery in the model was only calculated for the blocks that are either classified as a Measured or Indicated Resource. This is demonstrated by the relevant definitions for the CIM Standards/NI 43-101, which state that a Mineral Reserve is the economically mineable section of a Measured or Indicated Mineral Resource demonstrated by at least a Preliminary Feasibility Study. Using this definition, no recovery, and no economic value is given to the blocks within the model that are categorized as an Inferred Resource.

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15.1.2

Open Pit Optimization

In order to develop an optimal engineered pit design for the Rainy River deposit, an optimized pit shell was first prepared using the Lerchs-Grossman 3D routine in MineSight (LG 3D). The LG 3D pit optimizer algorithm will find a set of blocks with the maximum value per tonne, creating an optimized pit shell from the 3D block model.

With defined pit optimization parameters including gold and silver prices, mining, processing and other indirect costs, Au and Ag recoveries for each ore type (as determined from metallurgical testwork), pit slopes (as recommended by AMEC based on a geotechnical pit slope study) and other project related constraints, the pit optimizer searches for the pit shell with the highest undiscounted cash flow. In accordance with the guidelines of the NI 43-101 and the Canadian Institute of Mine Metallurgy and Petroleum Definition Standards for Mineral Resources and Mineral Reserves, only blocks classified as either Measured or Indicated are allowed to drive the pit optimizer for a feasibility study.

15.1.2.1

Pit Optimization Parameters

The main pit optimization parameters used in the LG 3D routine are listed in Table 15-2.

The costs for the pit optimization process were based on best available information at the time, including costs from the Preliminary Economic Assessment Update, some preliminary processing costs developed during the initial phase of the Feasibility Study, costs from similar mining operations and BBA experience.

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Table 15-2: Pit Optimization Parameters


Type of Activity	Unit	Values
Mining Cost¹	($/t mined)	1.89
Processing Cost¹	($/t milled)	8.73
General and Administration Cost¹	($/t milled)	1.00
Refining Cost¹	($/t milled)	1.50
Au Recovery	%	Varied as per equation
Ag Recovery	LOM average %	64.1
Gold Selling Price	USD$/oz.	Varied from 200 to 1250
Silver Selling Price	USD$/oz.	25
Exchange Rate	CAD/USD	1.05
Overall Pit Slope Angle	degree	Varied from 38 to 51
Overall Overburden Slope Angle	degree	16

^1.	Pit optimization parameters differ from the final operating cost figures.

The LG 3D pit optimization was run using complex slopes. It is important to note that all the slopes have been downgraded by 3 degrees on average from the final design specification provided by AMEC. This is done to account for operational design factors such as ramps, geotechnical berms, and benching arrangements which will be incorporated subsequently in the engineering design process.

Figure 15-2 shows the recommended complex slopes within the different zones (AMEC 2013F) of the deposit.

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LOGO

Figure 15-2: Slope by Sector for Pit Optimization

In addition to the aforementioned processing and slope parameters, there were also various limits and constraints that were imposed as agreed upon by BBA and Rainy River. These are as follows:

•

A constraint was provided from Bayfield’s property to the east of the pit (Rainy River owns surface rights to the Bayfield property. Mineral rights are owned by Bayfield Ventures Corp.);

•

Environmental and hydrogeological considerations require a 100 m buffer zone from the Pinweood River to the south of the pit; and

•

A -50 masl elevation constraint was applied to the pit optimization. In accordance with the PEA Update results, Rainy River Resources decided that the option for an open pit limited to a depth of -50 masl with the underground operation would be pursued for the Feasibility Study. The average topographic elevation of the pit is 350 m and is limited to a depth of 400 m.

The aforementioned constraints and buffer zones are indicated in Figure 15-3.

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LOGO

Figure 15-3: Surface Constraints for Pit Optimization

15.1.2.2

Pit Optimization Results

Using the technical and economic parameters described previously, the MineSight LG 3D pit optimizer tool was run and produced optimum pit shells at different gold prices for the Rainy River deposit. Once the series of pit shells were generated, total material moved, total in-pit resource and stripping ratios were evaluated to identify the optimal pit. Based on this analysis, the chosen optimized pit for this Feasibility Study was the pit having a gold price of USD $800/oz. This pit was chosen as it has one of the highest undiscounted cash flows and a mine life of approximately 15 years.

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A plan view of the LG 3D pit shell is shown in Figure 15-4. The pit edge commences at an elevation of 360 masl with the pit bottom is located at approximately – 50 masl.

The theoretical pit shell resulting from the LG 3D optimization is only preliminary and does not represent a practical design for mining. This optimized pit shell was used as a guide for the detailed mine design based on the required operational haulage ramp, proper pit slopes and benching arrangements as presented in Section 15.1.2.1.

LOGO

Figure 15-4: Rainy River Theoretical Pit Shell (Plan View)

15.1.2.3

Mill Cut-Off Grade

The break-even cut-off grade or the milling cut-off grade (COG) is used to classify the material inside the pit limits as rock or waste. For material located inside the pit, the break-even cut-off is

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the grade required to cover the costs for processing, G&A, and other costs related to gold refining and transport. The mill COG was calculated at 0.30 g/t Au, including an average dilution rate of 10%.

Table 15-3 presents the results of a sensitivity analysis of the undiluted Measured and Indicated Resources and gold grade versus cut-off grade (“COG”). As can be seen, Measured and Indicated Resources and Au grade show significant changes to COG variations between 0.2 and 0.4 g/t Au. The option with the lowest COG generating a positive cash flow is selected. This analysis confirms the selected COG of 0.30 g/t Au, and benchmarks well with similar gold operations.

Table 15-3: Selected Pits at Various Cut-Off Grades


Cut-Off Grade (g/t Au)	M+I Resources (Mt)		Undiluted Gold Grade (g/t Au)		Additional Gold (‘000 oz.)		Cash Flow¹ (M$)
0.40		84.1		1.25		—		—
0.45		94.9		1.15		119		32.0
0.30		108.7		1.04		130		16.6
0.25		126.3		0.93		140		-9.9
0.250		148.2		0.82		142		-51.3

¹	Cash flow based on a gold price of USD $1,250/oz., a processing cost of $8.65/t milled and a G&A cost of $1.21/t milled.

15.1.2.4

Equivalent Gold Grade

An equivalent gold grade (Au eq Grade) is used in order to take into account the silver revenues when using the gold COG to classify a block as ore or waste. The silver grade is transferred to its corresponding gold grade using the following calculation:

Equivalent Gold Grade Formula (Eq Au g/t)


		Eq Au (g/t) = Au (g/t) +		Ag (g/t) * Ag price ($/oz.) * Ag mill recovery (%)
				Au price ($/oz.) * Au mill recovery (%)

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15.1.3

Detailed Mine Design

The detailed mine design was carried out using the selected LG 3D pit shell as a guide. The proposed pit design includes the practical geometry required in a mine, including pit access and haulage ramp to all pit benches, pit slope design, benching configurations, smoothed pit walls and catch berms. The major design parameters used are described in Table 15-4.

Table 15-4: Detailed Open Pit Mine Design Parameters


Parameter		Value
Benching Arrangement		3 x 10 m
Berm Width		10.5 m
Inter-Ramp Angle (IRA)		40° - 56.3° (AMEC 2013F)
Bench Face Angle (BFA)		50° - 75° (AMEC 2013F)
Ramp Width (1-lane)		20 m
Ramp Width (2-lane)		33 m
Ramp Grade		10%

The in-pit haulage roads are 33 m wide to accommodate the proposed 226-tonne haul trucks. This ramp will provide enough room for 2-way traffic and minimize the truck cycle times. A single-lane ramp of 20 m wide will be placed near the bottom of the pit design in order to minimize the overall stripping ratio of the pit. All in-pit ramps have been restricted to a 10% gradient. All slope configurations and angles are based on recommendations indicated in the AMEC Geotechnical studies.

The in-pit haulage ramp exits on the east side of the pit in order to facilitate easy access to the site infrastructure, primary crusher, and to the waste rock piles. The designed pit is approximately 1,700 m in length by 1,450 m wide and 410 m deep. The lowest bench is at an elevation of 50 m below sea level.

Figure 15-5 presents a detailed plan view of the proposed open pit mine (final pit) and Figure 15-6 presents an isometric view of the final pit and the selected optimized pit shell. Figures 15-7 to 15-11 present typical bench plans and cross-sections of the detailed pit versus the optimized pit. A cut-off grade of 0.30 g/t Au eq is shown in the coloured blocks.

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LOGO

Figure 15-5: Detailed Open Pit Mine Design (Plan View)

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LOGO

Figure 15-6: Final Pit Design and LG Optimization – Isometric View

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LOGO

Figure 15-7: Final Pit Design and LG Optimization – Elevation 300 masl

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LOGO

Figure 15-8: Final Pit Design and LG Optimization – Elevation 150 masl

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LOGO

Figure 15-9: Final Pit Design and LG Optimization – Elevation 0 masl

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LOGO

Figure 15-10: Final Pit Design and LG Optimization – Cross Section (East 425 500, looking West)

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LOGO

Figure 15-11: Final Pit Design and LG Optimization – Cross Section (East 425 800, looking West)

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15.1.4

In-Pit Dilution and Mine Recovery

Using the resource block model, the dilution rate and the mine recovery were estimated for the mine. In the estimation, it was assumed that the selected mining method will be optimum, i.e., good blasting practice as well as a good practice of dilution control. Under this best case scenario, it was assumed that the main source of dilution and mine mineralized material loss will be at the contact between the mineralized material and waste using the following parameters:

•

The minimum mining width is 7 m (1 block);

•

Contact dilution of 1 m at the mineralized material/waste contact;

•

Contact mineralized material loss of 0.5 m when the mining width exceeds the minimum mining width (7 m); and

•

The “orphan” blocks are not mined.

The estimation of the in-pit dilution was carried out on five (5) selected equally spaced benches. The results are presented in Table 15-5.

Table 15-5: Estimation of In-pit Dilution and Mine Recovery


Parameter		Value
In-Pit Dilution		9.7%
Dilution Au Grade		0.22 g/t Au
Dilution Ag Grade		1.31 g/t Ag
Mine Recovery (ore loss)		95%(5%)

15.1.5

Open Pit Material Inventory

The open pit mine contains 113.2 Mt of diluted reserves in the Proven and Probable categories at an average grade of 0.97 g/t Au and 2.65 g/t Ag. Total waste material amounts to 350.6 Mt of waste rock and 80.0 Mt of overburden resulting in an overall open pit strip ratio of 3.8 (tonnes of waste rock and overburden per tonne of ore).

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15.2

Underground Mining

15.2.1

Underground Mineral Reserves

15.2.1.1

Resource Model

The Rainy River Gold Project mineral resources were provided to Golder by SRK in a CSV file format. This database was imported by Golder into Gemcom’s Surpac software for reserve calculations and mine design. The differences between the modelled values for tonnes, and gold and silver grades were determined to be within 1%.

15.2.1.2

Cut-off Grade Estimation

Initially, it was anticipated that the Rainy River Underground (“RRU”) would use both cut and fill (“CAF”) and longhole open stoping (“LHOS”) at a mine production rate of 2,000 tpd. Separate cut off grades (“COGs”) using gold equivalent (“Au eq”) were calculated for each mining method as the mining costs can vary significantly. The initial COG estimate for the LHOS was 2.0 g/t Au eq and for the CAF was 2.5 g/t Au eq. These estimates are based on Canadian mining industry average underground mine operating costs for the selected mining methods. The milling, general and administrative and selling costs, gold price, exchange rate and mill recovery inputs were provided to Golder by BBA.

In consultation with Rainy River Resources, it was decided to investigate the mining potential using a COG of 3.5 g/t Au eq for both CAF and LHOS at a reduced mining rate of 1,000 tpd. This option improves the margin per tonne of ore by requiring less mine development and capital cost, however, it also has a higher operating cost due to the decreased production rate. The reduced tonnage would increase the risk of there being insufficient resources underground to pay for the infrastructure and equipment required to mine it. The underground mine operating cost is discussed in Section 21 and is estimated to be $75.52 per tonne, which corresponds to a COG of approximately 2.5 g/t Au eq. Since the 3.5 g/t Au eq COG is higher than the estimated minimum required, and at this cut-off there is an underground mineral resource of approximately 4.9 Mt grading 6.40 g/t Au eq, it is believed that the 3.5 g/t Au eq option carries less risk. The underground design and cost estimate was based on the 3.5 g/t Au eq case.

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Within the 3.5 g/t Au eq design, a COG of 0.35 g/t Au eq was used to define incremental resource. This is material that is mined in the process of developing ore reserves (material above the 3.5 g/t Au eq), including sill drifts developed within mineralization but outside the mineral reserve envelopes and is economic to process once hauled to surface.

15.2.1.3

Mineral Resource Estimate

Using a COG of 3.5 g/t Au eq, there are 4.9 Mt grading 6.38 g/t Au and 6.58 g/t Ag available to the underground (Measured and Indicates reserves). These underground resources were evaluated outside of the feasibility pit design (Pit F), which was received as a RAR file from BBA on December 10, 2012. Table 15-6 contains a summary of the total and underground mineral resources for the RRGP. Refer to Section 14 for a description of the total mineral resource estimation by SRK.

Table 15-6: Combined Open Pit and Underground Mineral Resource Estimated by SRK and

Underground Mineral Resource Estimate


Category	Tonnes ‘000 t		Gold Grade g/t		Silver Grade g/t
Combined Open Pit and Underground (SRK February 24, 2012)^(a)
Measured		23,243		1.30		2.00
Indicated		127,337		1.14		2.54
Combined Measured and Indicated		150,580		1.17		2.46
Underground^(b)
Measured		280		6.72		3.41
Indicated		4,620		6.35		6.78
Underground Measured and Indicated		4,900		6.38		6.58

(a)	Open pit resources reported at a COG of 0.35 g/t Au and underground resources reported at a COG of 2.5 g/t Au.

(b)	Underground resources reported at a COG of 3.5 g/t Au outside the feasibility pit design.

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15.2.2

Underground Mineral Reserve Estimate

CAE’s Mineable Shape Optimizer (“MSO”) was used to estimate the in-stope mineral resources using the LHOS underground mining method. Multiple MSO evaluations were completed to determine the optimal sublevel spacing and hangingwall dip to recover the most resources. The final MSO input parameters used were ultimately governed by a desire to limit capital expense and are shown in Table 15-7.

Table 15-7: MSO Input Parameters


Parameter	Units	Min	Max
Cut-off grade	g/t	3.5
Stope dip	Degrees	120	120
Strike angle	Degrees	90	20
Waste pillar width	m	0.00001
Width	m	5	5
Height	m	25	25

The goal of MSO is to give engineers the ability to automatically create preliminary stope designs at a greatly improved speed. However, Golder’s experience indicates that the automatically designed stopes can sometimes produce unrealistic results in the form of irregular stope walls (hangingwall and/or footwall). To ensure a more realistic underground reserve estimate, each stoping area identified by MSO was manually smoothed, resulting in lower grade and higher tonnage. Approximately 80% of the underground reserves are contained in the LHOS stopes.

Longhole open stoping areas that do not form part of the reserve, or shallower dipping resources above the COG, were evaluated with the CAF mining method. The CAF stopes were designed using the resource model block dimensions (5 m wide by 5 m high by 5 m long). Simple economics were used to determine if an area should be added to an existing CAF lift. If a mining area was mineable by both CAF and LHOS mining methods, the method that either produced the highest margin per tonne of ore or that matched the methods used in surrounding areas was chosen. Approximately 20% of the underground reserves are in the CAF stopes.

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15.2.3

Dilution and Mine Recovery

The RRGP contains both longitudinal and transverse LHOS and CAF stopes. This section will discuss methods used to estimate the dilution from all sources and the mining recovery of each method. All dilution discussed in this section is considered unplanned dilution from stope wall failures, blast overbreak or backfill from stope walls and floors.

The assumptions presented in this section are based on the stope designs, as discussed in Section 16.3. However, definition drilling will be completed prior to developing and mining the proposed stopes. It is possible that the information gained during definition drilling could change the assumptions used to produce the dilution estimates.

Waste Rock Dilution

Waste rock dilution is mainly from material sloughing from the hangingwall with minor amounts from the stoping area ends. AMEC is the geotechnical consultant responsible for rock mechanics and have produced an Equivalent Linear Overbreak/Slough (“ELOS”) figure based on the stope design discussed in Section 16.3.1, showing the average ELOS values for the different longhole mining areas. It is estimated that, on average, the Rainy River Gold Project underground stopes will have between 0.25 m and 0.5 m of hangingwall slough.

To estimate the grade and tonnage of the ELOS, the stoping area shapes were expanded by 0.25 m in all directions and moved into the hangingwall, which created a 0.5 m skin along the hangingwall of the stopes and a 0.25 m skin at the stope area ends. This was completed for each stoping area, and these expanded shapes were included in the production schedule on which the mineral reserves are based. Overall, it is estimated that there will be approximately 6% dilution from the hangingwall and stope ends with an average grade of 2.66 g/t Au and 2.18 g/t Ag (Table 15-8).

Development of the CAF headings will require strict geological control to follow the mineralization and ensure minimum waste inclusion. At this time, it is understood that the gold grade is not controlled by a well-defined vein, which will make following the mineralization difficult. Assuming that the mineralized envelope (the lift length and width) is properly defined, the main source of

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waste rock dilution will come from sloughing of either wall of the CAF drift heading. It is estimated that the walls will contribute 4% dilution with an average grade of 1.37 g/t Au and 9.36 g/t Ag (Table 15-8). The majority of the CAF tonnage is located in the 17 east mining area which has higher silver grades.

Table 15-8: Total Dilution Estimates from Waste Rock by Mining Method


Mining Method	Average Depth (m)		Percentage			Tonnage (kt)		Au Grade (g/t)		Ag Grade (g/t)
LHOS		0.5		5.7	%		141.0		2.66		2.18
CAF		0.2		4.0	%		25.0		1.37		9.36
Total		—		5.4	%^(a)		166.0		2.47		3.26

(a)	This is the average percent dilution from waste rock for the entire underground mineral reserves.

Backfill Dilution

Backfill dilution at the RRGP will come from two (2) sources; mucking the floors in the LHOS and the CAF stopes, and sloughing of cemented rock fill (“CRF”) walls in the secondary and longitudinal LHOS. In the LHOS, the majority of mucking will be done by remote methods which will result in less operator control and a greater occurrence of digging into the floor. It is estimated that this form of dilution will amount to an average of 0.5 m from the floor of each stope panel.

Similarly, in the CAF cuts, backfill dilution will occur as a result of mucking from rock fill floors along the full length of the cuts. It is estimated that the average depth of over-digging the floor will be 0.25 m. This form of dilution can be better-controlled than in LHOS since all mucking is done conventionally i.e., not remotely. A summary of the dilution expected from mucking backfill is shown in Table 15-9.

Table 15-9: Total Dilution Estimate from Mucking Backfill by Mining Method


Mining Method	Depth (m)		Percent (%)			Tonnage (kt)
LHOS		0.5		2.0			49.7
CAF		0.25		5.0			31.3
Total		—		2.6	^(a)		81.0

(a)	This is the average percent dilution from mucking backfill for the entire underground mineral reserve.

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Finally, additional backfill dilution is expected in the secondary stopes from CRF sloughing from the side walls and in longitudinal stopes from the CRF stope end wall. It is estimated that an average of 0.25 m of backfill dilution will result from the CRF walls. A summary of the dilution expected from CRF by LHOS mining method is presented in Table 15-10.

Table 15-10 Total Dilution Estimate from Cemented Rock Fill Wall Sloughing by Mining Method


Mining Method	Depth (m)		Percent (%)			Tonnage (kt)
Secondary LHOS		0.3		3.0			41.8
Longitudinal LHOS		0.6		2.8			7.8
Total		—		1.6	^(a)		49.6

(a)	This is the average percent dilution from cemented rockfill wall sloughing for the entire underground mineral reserve.

Dilution Summary

Table 15-11 below summarizes the estimated quantity and grade of the dilution for each mining method.

Table 15-11: Total Dilution Estimate by Mining Method


Mining Method	Percent (%)			Tonnage (kt)		Gold Grade (g/t)		Silver Grade (g/t)
LHOS		9.7			240.0		1.56		1.28
CAF		9.0			56.3		0.61		4.16
Total		9.6	^(a)		296.3		1.38		1.83

(a)	This is the average percent dilution from all sources for the entire reserve.

15.2.4

Mining Recovery

Mining recovery factors are applied to each mining method to account for material that is part of the stope design but is not recovered. These losses can occur from inefficient drilling and blasting in stope corners and walls, difficult remote mucking in stope corners and edges, or abandoning a stope due to excessive dilution from waste rock or CRF. Due to the shallow dip of the hangingwall, the recovery in the longitudinal areas might be lower than indicated. Conversely, the average stope at the RRGP is small, and losses in stope corners could be mitigated by efficient stope mucking.

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Since the longitudinal areas make up 10% of all longhole areas, it is felt that these reasons offset each other and that the same recovery factor can be applied to all LHOS areas. The factors used to estimate the mineral reserve are shown in Table 15-12.

Table 15-12: Mining Recovery Factors used to Estimate the Mineral Reserves


Mining Method	Recovery Factor
LHOS		95	%
CAF		95	%

15.2.5

Underground Material Inventory

The mineral reserves are calculated using a gold price of $1,400 per ounce, a silver price of $25 per ounce and a discount rate of 5% as provided by Rainy River Resources.

According to 43-101 standards, Measured resources are typically converted into the proven reserves category, and Indicated resources are converted into probable reserves. In discussion with Rainy River and SRK, it was decided to classify all the underground reserves as probable, regardless of their category. The Measured resources were defined with open pit extraction in mind and a strict criterion was assigned to define Measured resources. The only Measured underground resources at Rainy River are those immediately adjacent to those defined as Measured for open pit extraction. To assign the highest confidence category to the underground resources will require some additional evidence of grade continuity, such as underground channel sampling. The mineral reserves for the RRU are presented in Table 15-13 by mining method, and in Table 15-14 by reserve category. Approximately 63% of tonnes and 49% of gold ounces have been converted from mineral resources to mineral reserves for the RRU.

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Table 15-13: Total Underground Mineral Reserves by Mining Method (COG of 3.5 g/t Au eq)


Mining Method	Tonnage ‘000 t		Gold Grade g/t		Silver Grade g/t		Contained Gold ‘000 oz.		Contained Silver ‘000 oz.
Development Reserves		95		5.44		3.06		17		9
LHOS Reserves		2,364		5.29		3.04		402		231
CAF Reserves		648		4.22		20.56		88		428

Table 15-14: Total Underground Mineral Reserves by Category (COG of 3.5 g/t Au eq)


Underground Mineral Reserves	Tonnage ‘000 t		Gold Grade g/t		Silver Grade g/t		Contained Gold ‘000 oz.		Contained Silver ‘000 oz.
Probable		3,107		5.07		6.69		507		668

15.3

Open Pit and Underground Mineral Reserves

In accordance with the NI 43-101 standards of mineral classification, the measured and indicated resources inside the final pit limits have been transferred into Proven and Probable reserves. All underground reserves have been transferred into Probable reserves.

Table 15-15 shows a detailed summary of the in-pit and underground reserves.

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Table 15-15: Open Pit and Underground Proven and Probable Mineral Reserves (April 10, 2013)^1,2,3


Reserves	Tonnage		Au Grade		Ag Grade		Au		Ag
Category	Mt		g/t		g/t		In-Situ oz.		In-Situ oz.
Open Pit
Proven		27.7		1.14		1.94		1,014,584		1,727,979
Probable		85.5		0.91		2.88		2,510,641		7,918,793

TOTAL		113.2		0.97		2.65		3,525,225		9,646,772

Underground
Proven
Probable		3.1		5.07		6.69		506,283		668,240

TOTAL		3.1		5.07		6.69		506,283		668,240


Combined
Proven		27.7		1.14		1.94		1,014,584		1,727,979
Probable		88.6		1.06		3.01		3,016,924		8,587,034

TOTAL		116.3		1.08		2.76		4,031,508		10,315,013

Qualified persons: The mineral reserve statement was prepared by Patrice Live (OIQ #38991) of BBA and Donald Tolfree (APEGBC #32557), of Golder, both “independent qualified persons” as that term is defined in National Instrument 43-101. Rainy River’s engineering assessment in Richardson Township is being supervised by Garett Macdonald, P.Eng. (PEO #90475344), Vice-President, Operations and a Qualified Person as defined by National Instrument 43-101. Mineral resource estimates may be materially affected by environmental, permitting, legal, title, taxation, sociopolitical, marketing and other relevant issues.

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16.

MINING METHODS

16.1

Introduction

The Feasibility Study assumes both open pit and underground mining methods will be used for resource extraction. The mining methods and production capacity have been chosen to match a milling throughput rate of 21,000 tpd (20,000 tpd from the open pit and 1,000 tpd from underground when full production is achieved). The location of the open pit and the underground access ramp (main portal) is shown on the project general layout in Appendix H. Both the open pit and underground operations will deliver material to a common gyratory crusher for primary size reduction and delivery to the processing plant. Utilization of Rainy River’s mining equipment and personnel is envisaged for the development of the open pit as well as for the removal of overburden. Underground contractors will be used for initial development of the underground mine, followed by RR development and production crews during mine operations.

Figure 16-1 shows a representation of the underground and open pit designs.

LOGO

Figure 16-1: Isometric View of the Rainy River Underground Mine

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16.2

Open Pit Mining Methods

16.2.1

Open Pit and Stockpiles Geotechnical Designs

16.2.1.1

Open Pit Geotechnical Design

The feasibility level site investigation work and open pit slope design criteria were developed by AMEC (2012F; 2012G, 2013D, 2013E, 2013F), and provided to BBA for open pit development. The approximate dimensions of the pit will be 1,600 m [length] x 1,600 m [width] x 400 [depth] (in the main south pit, ODM Zone). The northern section of the pit (433 Zone) will be approximately 325 m deep.

The bedrock geology is composed of metavolcanic rock that hosts the multi-lens gold deposit. These lenses dip between 50° to 65°. The bedrock for the most part is overlain by 10 to 40 m of overburden consisting of primarily silty clays separated by clay till, with gravelly sand till and boulders directly overlaying the bedrock. The assessment of various rock structural domains was based on the analysis of 10 NQ-sized boreholes, with geomechanical logging of oriented core and packer testing. These boreholes were oriented in various azimuths, dipping at around 65° and totalling 4.4 km of drilling. Additionally, ten (10) exploration boreholes were surveyed by DGI Geoscience Inc. using acoustic and optical televiewer tools.

The primary jointing identified at the site is the foliation set, which follows the dip of the ore lenses, strikes approximately east-west and dips to the south, at approximately 55° (in the south ODM pit) or 49° (in the north 433 pit). The two (2) other sets are sub-vertical, striking essentially north-south, and in the south pit dip towards the west at approximately 75°, while in the north pit dips towards the west at 85°. A persistent east dipping fault striking north south transects the pit, however, due to the fault orientation and only a localized increase in the fracture frequency, the fault is not expected to impact the overall wall stability. The rock mass typically observed has two (2) to two plus (2+) random major joints sets with a “good” rock mass rating and geological strength index (“GSI”) ranging from 64 to 72. A total of 176 core samples were collected for strength testing with 268 test specimens prepared. A total of 117 specimens were tested for uniaxial compressive strength (“UCS”); 42 for triaxial strength; 100 for Brazilian tensile strength, and nine (9) for direct shear tests of open joints (AMEC, 2013D). Based on laboratory strength testing, felsic volcanics

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(the predominant wall rock) were found to be strong, to very strong rocks ranging in UCS from 71 to 116 MPa.

Design sectors for the open pit were based on probabilistic kinematic analysis to identify plane, wedge and toppling failure cumulative probabilities; plane and wedge failure probabilistic stability analyses; as well as probabilistic overall slope stability analyses using limit equilibrium (Figure 16-2). Most bench face angles (“BFAs”) will be cut at 70° to 75°, with inter-ramp angles (“IRAs”) comprised between 54.5° and 56.3° and overall slope angles (“OSAs”) in bedrock of 41 to 45, with an average bench width of 10.5 m and final bench height of 30 m. To ensure stability, bench widths will be increased to 12 m to prevent toppling failure in the south walls of both pits as well as in the west wall of the south pit. Additionally, BFAs will be reduced to 50° or 55°, to limit planar or wedge failures in the north walls of both pits. A summary of the design guidelines and bench configurations can be seen in Figure 16-2 and Table 16-1. A 20 m set-back at the overburden-bedrock interface will be placed to allow access and monitoring of overburden slopes and the pit. A 20 m geotechnical berm is recommended for the south wall of ODM, in which the inter ramp spacing is greater than 200 m. Under partially saturated conditions, with a disturbance factor of 0.7 and a seismic (pseudo static) load of 0.1, the worst case scenario for the south wall of ODM will have a deterministic factor of safety (“FOS”) of 1.29, and a probability of failure (“POF”) of 7.7 %, which meets the accepted minimum FOS used for designing operating open pits. The other bedrock walls typically have FOS ranging from 1.37 to 2.5 for similar conditions (AMEC, 2013F).

Table 16-1: Recommended Overall Slope Geometry by Sector (AMEC, 2013F)


Rock	Segment	Bench Face Angle (^o)		Bench Height (m)		Bench Width (m)		IRA (^o)
FLS	5, 10		50		3 x 10		10.5		40.0
FLS	3		55		3 x 10		10.5		43.6
FLS	4, 6, 7, 9		70		3 x 10		10.5		54.5
FLS	1		73		3 x 10		12		54.8
FLS	2, 8		75		3 x 10		12		56.3

Note:

Segments 3 t and 7 t (Figure 16-2) should be transition zones which require that the BFA progressively change from 55° to 70° both from west to east to ensure stability of the walls.

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In order to maintain short and long-term stability of the overburden material, the slopes in the overburden will range from 20° (2.75W:1H) to 17° (3.25W:1H) for overburden less than 25 m thick, and between 17° (3.25W:1H) to 14° (4W:1H) for overburden greater than 25 m, respectively (AMEC, 2012F).

LOGO

Figure 16-2: Open Pit Design Zones & Recommendations (AMEC, 2013F)

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16.2.1.2

Waste Rock and Overburden Pile Design

Geotechnical slope stability recommendations were provided by AMEC (2013c) for the mine waste rock and overburden stockpiles. The rate of stockpile development will control stability due to excess pore pressures generated in the foundation soils, as described in AMEC (2013a). From a stability perspective, the slope at the south perimeter of the West Stockpile and west side of the East Stockpile are critical as the height is greatest and the upper varved clay strata the thickest.

Mine rock stockpile slopes of 6H:1V have been adopted in the critical areas. The high in-situ moisture contents of the end-dumped overburden, which will have low strength and tend to develop high induced pore water pressure during construction, dictate shallower external slopes of 8H:1V to meet the required factors of safety.

Stockpile construction will follow the observational approach. If higher than expected pore pressures or significant deformations in the foundation are noted, then mitigative measures such as flattening the overall slope or raising the toe berm along the south side of the West stockpile will be implemented to ensure stability, in particular adjacent to the Pinewood River.

16.2.2

Open Pit Mine Planning

16.2.2.1

Whittle Consulting Enterprise Optimization (Whittle, 2012)

In July 2012, Ausenco (in partnership with Whittle Consulting) provided recommendations on the open pit mine planning strategy based on the Whittle Consulting Enterprise Optimization (“EO”) methodology; an in-house strategic tool for mining operations. EO methodology is an integrated approach to maximizing the NPV of a mining operation by simultaneously optimizing 10 different mechanisms across the mining value chain. The main goal is to isolate the critical cost drivers and maximize value throughout the mining system. Contained in the business model are a series of calculations that are performed on every block in the resource model to generate revenue and cost fields that are then used for pit optimization and scheduling by Prober.

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The EO Study was conducted to support the PEA Update and subsequent Feasibility Study for the Rainy River Gold Project. The EO process is iterative and the business model tends to evolve with the costs and resource model. Two (2) generations of resource models have been used thus far in the Study, with several fixed gold recoveries. Costs also evolved slightly during the PEA Update process.

Analyses conducted for the PEA Update and FS, using June 2011 block model, include:

•

Base case assessment;

•

Variable cut-off grade and stockpile value contributions;

•

Mining capacity trade-off;

•

Processing capacity trade-off;

•

Underground cut-over trade-off,

•

Pit and phase design based on Enterprise Optimization economics using the theory of constraints and activity-based costing; and

•

NPV-optimized schedules calculated by Whittle Consulting’s proprietary software, Prober.

Whittle Consulting generated a series of mining scenarios by varying different parameters such as mining capacity, processing capacity and depth limit. BBA was provided a pit phasing and stockpile movement strategy by period which was used as a guideline for detailed mine planning. The results of the mining rate assessment provided in the EO study suggested that a constant mining rate of 100 Mtpa of total material moved would yield the highest NPV. Since the initial EO study was based on the PEA results for a milling rate of 32 ktpd, the mining rate was adjusted to approximately 60 Mtpa to match the selected milling rate of 21 ktpd in the current Feasibility Study. The use of an elevated cut-off grade strategy over the life of mine has resulted in an large amount of stockpiled material .

16.2.3

Open Pit Mine Production Schedule

An open pit mine production schedule was prepared by BBA for the development and the operation of the Rainy River Gold Project. The schedule is shown in Table 16-30. The mining production schedule is based on a pre-stripping period of 24 months, starting in Q3 of 2014 and ending in Q3 of 2016, and mill start-up beginning in Q3 of 2016. Based on the total in-pit reserves

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available, the open pit mine life is 16 years using a production rate ranging from 20,000 to 21,000 tpd.

In order to optimize the operational stripping ratio in the early years of the Project, and to increase the net present value of the Project, an optimized shell for a starter pit, representing approximately five (5) years of mining, was generated using the LG 3D MineSight routine. This optimized starter pit shell is presented in Figure 16-3 and was used as a guide to prepare the first few years of the LOM mine plan. Based on a strategic mine plan study prepared by Whittle Consulting, with the objective to maximize the Project’s IRR and NPV, an elevated COG was strategically employed from years 2016 to 2024. Material between the LOM COG (0.30 g/t Au eq) and the yearly COG was stockpiled for processing at the end of the mine life. The total material stockpiled during the open pit operations amounts to 43.1 Mt at an average grade of 0.37 g/t Au and 1.97 g/t Ag. This material is reclaimed at the end of the open pit life, or, as needed to ensure mill throughput is maximized at the end of the mine operation.

LOGO

Figure 16-3: Starter Pit and Final Pit – 3D View

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The actual open pit pre-production and production periods total two (2) and ten (10) years, respectively, and the low-grade stockpile reclaim period is approximately six (6) years in duration. The total combined remaining-life-of-mine (RLOM) ore and waste pre-production quantity is approximately 11 Mt in 2014, and ramps up to a maximum annual production rate of approximately 62 Mt in 2017 to 2023. Mine production slowly decreases until the pit is depleted at the beginning of 2026.

The mill start-up assumes a ramp up of 25% the first month, 50% the second month, 75% the third month and full production the fourth month. The mill will be fed with ore from the open pit at a rate of 21,000 tpd until 2018. During the period between 2019 to 2028, the open pit mine and underground mine will feed approximately 20,000 tpd and 1,000 tpd of ore, respectively, to the primary crusher.

The open pit production schedule has been developed on a quarterly basis for the first four (4) years and annually for the rest of the mine life. Figure 16-4 to Figure 16-8 show a series of key end-of-period maps.

LOGO

Figure 16-4: Open Pit Mine Planning – 2016 Q2 (End of Pre-Production Period)

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Figure 16-4 presents the contours of the Phase 1 (initial) pit at the end of the pre-production period. During Years -1 and -2, 30.7 Mt of overburden and 21.8 Mt of rock will be excavated from the initial pit to provide access to the ore for the beginning production in Q3 of 2016 (Year 1).

LOGO

Figure 16-5: Open Pit Mine Planning – 2017 (End of Year 2)

The final pit starts stripping in Year 2. The contour at the end of that year is shown in Figure 16-5.

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LOGO

Figure 16-6: Open Pit Mine Planning – 2019 (End of Year 4)

Figure 16-6 shows the pit at the end of Year 4. At this point, the production of ore in the final pit has started.

LOGO

Figure 16-7: Open Pit Mine Planning – 2021 (End of Year 6)

Figure 16-7 shows the pit at the end of the initial pit ore production in Year 6.

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LOGO

Figure 16-8: Open Pit Mine Planning – 2023 (End of Year 8)

Figure 16-8 shows the progress of the final pit at the end of Year 8. The final pit continues to be mined until its depletion at the beginning of Year 11.

16.2.3.1

Stockpile Re-handling

During the open pit operation, 43.1 Mt of low-grade ore at an average grade of 0.37 g/t Au and 1.97 g/t Ag are stockpiled in an effort to increase the feed grade. The low-grade ore stockpile is located between the PAG pile and the primary crusher, in the east dump area. The east dump is approximately 1 km (road distance) from the primary crusher. After the pit is depleted, the mill is fed at a rate of 20,000 to 21,000 tpd from the low-grade ore stockpile from 2026 to 2031, depending on the underground production in years 2026 to 2028.

The open pit material movement trends over the life of the mine can be seen graphically in Figure 16-9.

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Figure 16-9: Open Pit Material Movement over the Life-of-Mine

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16.2.4

Material Management

During the pre-production and production stages of the Project, a significant amount of waste rock and overburden material is removed and stored in nearby disposal areas. A portion of this material will be used for tailings dam and road construction as well as for reclamation of the site. The waste rock and overburden piles presented in this section were designed without deducting any waste material that might be used for site construction. All design parameters for the various stockpiles are based on the AMEC’s recommendations (AMEC 2013D, E, F, G).

The waste rock material will be placed onto two (2) waste piles: the non-potentially acid generating pile (“NPAG”) and the potentially acid generating pile (“PAG”). The neutralization potential ratio (“NPR”) is used to classify the waste rock: the waste rock with a NPR > 2 is considered as NPAG and the waste with a NPR £ 2 is considered as PAG. According to the NPR, approximately 40% of the waste rock in the open pit is considered to be PAG material.

There are two (2) areas for material disposal - the West Dump and the East Dump. The West Dump area is located west of the pit and contains the NPAG and the overburden material (“OB”). The East Dump area is located east of the pit contains the PAG and low grade ore material. The general arrangement of the piles around the pit can be seen in Figure 16-10.

Plan views of the stockpiles with slope and height details are presented in Figure 16-11 and Figure 16-12.

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LOGO

Figure 16-10: Site Plan Showing West and East Dump Areas

LOGO

Figure 16-11: East Dump Area

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LOGO

Figure 16-12: West Dump Area

16.2.4.1

Waste Rock Pile Design

The waste rock piles have been designed according to the waste requirements of the pit and are located around the periphery of the mine to minimize the haulage distance. The material properties assumptions used for the design of the waste rock piles are an in-situ waste rock density of 2.8 t/m³ and a swell factor of 30%.

The NPAG and PAG piles are located and sized to fit entirely within Rainy River’s mining claims or leasing claims and are kept at an adequate distance from all major water basins.

The waste rock stockpiles were designed according to the design parameters presented in Table 16-2. These parameters were selected according to an inter-ramp slope of 6H:1V, as recommended by AMEC. This inter-ramp slope has been used for the external stockpile slopes only. A steeper angle of 3H:1V has been used for internal slopes as shown in Figure 16-11 and Figure 16-12. Due to the bearing capacity of the ground in the south area of the East dump, AMEC

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also recommends that a shallower inter-ramp angle of 8H:1V be used for the south slope of the east dump.

Table 16-2: Waste Rock and Low-Grade Ore Stockpile Design Parameters


Waste Rock and Low-Grade Ore Piles	Value	Unit
Bench Face Angle	33	degrees
Inter-ramp Slope Angle	Varied from 3H:1V to 8H:1V	degrees
Catch Bench Width	Varied from 20 to 66	m
Bench Height	15	m
Ramp Width	33	m
Ramp Grade	10	%
Swell Factor	30	%

The NPAG and PAG dumps have a capacity of 95.4 Mm³ and 67.4 Mm³, respectively. The dumps will be built in 5 m lifts, with 15 m bench heights. Dumping has been sequenced in phases to allow for shorter hauls during earlier years of operation. The design summary of the NPAG and PAG piles can be found in Table 16-3 and Table 16-4.

Table 16-3: NPAG Pile Summary


NPAG Pile	Value		Unit
Height		60	m
Top Elevation		410	masl
Footprint Area		2.1	M m²

Table 16-4: PAG Pile Summary


PAG Pile	Value		Unit
Height		30	m
Top Elevation		407	masl
Footprint Area		2.9	M m²

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16.2.4.2

Low Grade Ore Stockpile Design

The low-grade ore material between the LOM COG and the yearly COG is stockpiled in a reserved area of the east dump beside the PAG material. It is located close to the crusher to reduce the haulage distance during the reclaiming process. The material property assumptions used for the design of the low-grade ore stockpile are an in-situ ore density of 2.85 t/m³ and a swell factor of 30%. The pile is located entirely within Rainy River’s mining claims. The low-grade ore stockpile has a capacity of 20.6 Mm³ and is designed according to the same design parameters as the waste rock piles in Table 16-5. The design summary of the low-grade ore stockpile can be found in Table 16-7.

Table 16-5: Low-Grade Ore Stockpile Summary


Low-Grade Stockpile	Value		Unit
Height		45	m
Top Elevation		422	masl
Footprint Area		0.64	M m²

16.2.4.3

Overburden Pile Design

Overburden material will be removed during the period of Year -2 to Year 5. The overburden pile, located west of the pit, will have a capacity of 53.3 Mm³. The material properties assumptions used for the design of the overburden piles are an in situ density of 1.8 t/m³ and a swell factor of 20%. The pile is entirely located within Rainy River’s mining claims or leasing claims.

The overburden pile is designed according to the design parameters presented in Table 16-6. These parameters were selected according to an inter-ramp slope of 8H:1V, as recommended by AMEC (AMEC 2013).

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Table 16-6: Overburden Pile Design Parameters


Overburden Pile	Value	Unit
Bench Face Angle	20	degrees
Inter-ramp Slope Angle	8H:1V
Berm Width	81	m
Bench Height	15	m
Ramp Width	33	m
Ramp Grade	10	%
Swell Factor	20	%

The dump will be built in 5 m lifts, with 15 m bench heights. The design summary can be found in Table 16-7.

Table 16-7: Overburden Pile Design Summary


Overburden Pile	Value		Unit
Height		60	m
Top Elevation		410	masl
Footprint Area		2.03	M m²

16.2.4.4

In-Pit Dumping

In order to reduce the cycle time and the quantity of haul trucks and to limit the environmental impact of the stockpile, an in-pit dumping area will be used for NPAG material. The dump is located on the north extension of the pit and has a capacity of 15 Mt of NPAG waste rock. The area will be used as an in-pit dump once the 433 ramp has been fully developed and the 433 exhaust raise can be accessed from the main ramp. Figure 16-13 shows an isometric view of the in-pit dump.

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LOGO

Figure 16-13: In-Pit Dumping Area

16.2.5

Open Pit Mine Equipment and Operations

The Rainy River deposit will be mined using conventional open pit mining methods based on a truck/shovel operation. All equipment will be diesel-powered except for two (2) hydraulic shovels. Using the production schedule presented in Table 16-30, the mining fleet requirement was calculated. All equipment is assumed to be owned, operated and maintained by Rainy River.

Open pit mine operations are based on 720 shifts per year, and corresponds to operations running 2 x 12 hour shifts per day, 7 days per week and 360 days per year, with the assumption that five (5) operating days will be lost on average due to weather. The mining operations division will consist of the pit operations, maintenance, engineering and geology departments.

The mining methods are based on conventional drilling and blasting followed by loading and hauling. The selection of the primary fleet is based on operating time assumptions, mechanical availability and utilization, haulage distance and cycle time assumptions, and truck speed and fuel consumption profiles. The primary mining fleet consists of the following:

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•

The primary loading equipment for overburden, waste rock and ore are: one (1) diesel hydraulic shovel with a rated bucket capacity of 28 m³and two (2) electric hydraulic shovels with a rated bucket capacity of 28 m³. One (1) wheel loader (15 m³class) will be used on an as needed basis to complete the loading equipment fleet. The flexibility of the loader, with its fast response time, justifies its use in replacing a shovel in loading support activities;

•

The haul truck fleet is based on trucks with a 226-tonne payload, which is a good match with the 28 m³hydraulic shovels. The haul truck fleet commences with six (6) trucks in the pre-production phase and will increase to 19 trucks in Year 2018 (Year 3 of production); and

•

Production drilling will be accomplished using a fleet of diesel powered DTH blast hole rigs drilling 8¹/₂” diameter holes.

16.2.5.1

Operating Time Assumptions

The productive operating time available for each shift has been calculated for two (2) categories: 1) primary equipment; and 2) drills. They are separated in order to take into account extra scheduled delays typically associated to the drills, such as additional time required for moving between drill patterns and spotting time between blast holes.

Scheduled delays for the primary equipment and drills take into account operator lunch breaks, inspection and fueling, shift changes, and coffee breaks. Table 16-8 provides information about the scheduled delays considered.

Table 16-9 shows how net operating hours (“NOH”) are derived from scheduled delays, unscheduled delays and the job efficiency factor (“JEF”). Unscheduled delays are defined as delays that are outside of human control and those which cannot be predicted or planned. These types of delays can take into account traffic delays, matching factors and the efficiency of equipment movement. These factors were estimated based on similar operations.

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Table 16-8: Operating Shift Parameters


Shift Parameters	Value		Unit
Shift Day		2
Worker and Equipment Shift Operating Time
Shift Change		15	min
Inspection		15	min
Coffee Break		20	min
Lunch Break		30	min
Job Efficiency Factor (JEF)		83	%
Drills Operating Time
Shift Change		15	min
Inspection		15	min
Coffee Break		20	min
Lunch Break		30	min
Job Efficiency Factor (JEF)		75	%

Table 16-9: Equipment and Worker Operating Time


Operating Time Calculations	Value		Unit
Worker and Equipment Operating Time
Scheduled Time		720	min
Scheduled Delays		80	min
Scheduled Operating Time		640	min
Unscheduled Delays		107	min
Total Delays		187	min
Net Operating Time		533	min
Net Operating Hours		8.89	hr
Drills Operating Time
Scheduled Time		720	min
Scheduled Delays		80	min
Scheduled Operating Time		640	min
Unscheduled Delays		160	min
Total Delays		240	min
Net Operating Time		480	min
Net Operating Hours		8.00	hr

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16.2.5.2

Equipment Availability and Utilization

For each piece of major equipment, mechanical availability and utilization factors were designated. The mechanical availability is a percentage that represents the hours when the equipment cannot be operated due to planned maintenance or breakdowns (unplanned). These factors were derived from supplier recommendations and/or experience. Equipment utilization, also referred to as the “use of availability”, refers to the time that a piece of equipment is available and operated productively. The availability and utilization factors used over the LOM are illustrated in Table 16-10.

Table 16-10: Major Equipment Availability and Utilization


	Years
Equipment	2014			2015			2016			2017			2018			2019 - 2026
Haul Trucks
Haul Truck Availability		90	%		89.5	%		89.1	%		88.8	%		88.2	%	87.5% – 88.5%
Haul Truck Utilization		90	%		90	%		95	%		95	%		95	%	95%
Shovels
Diesel Shovel Availability		88.3	%		87.4	%		84.7	%		81.7	%		83	%	83% – 85%
Diesel Shovel Utilization		95	%		95	%		95	%		95	%		95	%	95%
Electric Shovel Availability		89	%		89	%		88.5	%		87.8	%		86.3	%	83% – 85%
Electric Shovel Utilization		95	%		95	%		95	%		95	%		95	%	95%
Drills
Drill Availability		83	%		83	%		83	%		83	%		83	%	83%
Drill Utilization		95	%		95	%		95	%		95	%		95	%	95%

16.2.5.3

Drilling and Blasting

The drill and blast design for the Study was determined by BBA, in collaboration with explosive suppliers familiar with this type of operation.

The ore zones (“OZ”) will be drilled using 8-¹/₂ inch diameter holes on a drilling pattern of 5.5 m spacing x 6.5 m burden. Waste rock areas will use the same hole diameter, but a slightly larger

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drilling pattern of 6.5 m x 7.0 m. The spacing and burden for the ore zone is tighter to produce better fragmentation and selectivity.

Holes will be drilled to a total depth of 11.2 m, including 1.2 m of sub-drilling for a 10 m bench height. A stemming height of approximately 4.0 m will be used to maximize the explosive column’s effectiveness. Based on the production schedule, up to three (3) drills will be required.

Blasting will be executed under contract with an explosive company that will supply the blasting materials and technology, as well as the equipment to store and deliver the explosive products. The explosives will be manufactured on-site by the explosives supplier in a purpose built bulk emulsion plant. The explosive supplier will also be responsible for providing a down-the-hole service. In order to obtain optimum fragmentation, which will improve the materials handling operations, high-precision electronic detonators will be used.

Blasting will be conducted using a 100% emulsion-type explosive with an average density of 1.25 kg/m³. A bulk emulsion was selected as it is easily transportable and has a lower environmental impact than other types of explosives resulting in lower ammonium nitrate levels emitted into the watershed.

Based on the drilling pattern described above, the powder factor has been estimated at 0.32 kg/tonne in ore and 0.26 kg/tonne in waste.

It is also assumed that pre-split will be required for final walls (inital pit and final pit).

A summary of the drill and blast specifications can be found in Table 16-11.

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Table 16-11: Drill and Blast Specifications


Drill Specifications
Parameter	Unit	Ore		Waste
Hole diameter	mm		215.9		215.9
Hole area	m²		0.0366		0.0366
Bench height	m		10		10
Sub-drill	m		1.2		1.2
Stemming	m		4.0		4.0
Loaded length	m		7.2		7.2
Hole spacing	m		5.5		6.5
Burden	m		6.5		7.0
Penetration rate	m/hr		40.0		40.0
Re-drill	%		10		10
Rock mass/hole	t		1019		1274

Bulk Emulsion
Density	kg/m³		1250		1250
kg/hole	kg/hole		329		329
Powder factor	kg/tonne		0.323		0.259

16.2.5.4

Loading and Hauling

Production will be carried out using a fleet of 226-tonne capacity dump trucks and hydraulic shovels with a bucket capacity of 28 m³ in ore and waste rock. The number of trucks operating at any given time is dependent on the annual production rate, and varies over the mine life. This fleet combination should allow for 4 pass-loading of trucks hauling ore and waste and 5 to 6 pass-loading of trucks hauling overburden. A maximum of 19 haul trucks will be required in the peak years.

The maximum shovel productivity per shift has been estimated at 36,500 tonnes of ore or waste per shift and 23,000 tonnes of overburden. Loading operations will also be assisted by one (1) large wheel loader to maximize the flexibility of the operation. The loader will be used as

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production equipment but also as a replacement for the shovel in down-time situations, as well as for other tasks involving material handling.

Average annual haul profiles were created for ore, waste rock and overburden. The haulage distances were further divided for in-pit flat hauls, in-pit ramp hauls, flat on topography hauls and for crusher and waste piles. In the MineSight software, haul routes were traced according to mining centroids for every bench (and material) for each year. Subsequently, with these centroid distances and the respective tonnage per bench (per material) mined, the weighted and averaged distances were calculated on a yearly basis. The in-pit ramp distances were also averaged in the same manner.

In order to optimize the waste cycle times for operation, dumping has been sequenced in phases to allocate shorter hauls during earlier years of the LOM. Centroid and up-ramp distances were traced for the waste pile locations and crusher location.

Haulage travel speeds and fuel consumptions for the trucks were based on supplier experience and were fine-tuned using factors from BBA’s equipment database. The travel speeds and fuel consumptions are shown segmented by type of haul in Table 16-12.

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Table 16-12: Trucks Speeds and Fuel Consumptions (Loaded and Empty)


	Haul Truck Loaded
Parameter	Acceleration 100 m		Flat (0%) Topo		Flat (0%) In-Pit/ Crusher/ Dump		Slope Up (10%)		Slope Down (-10%)		Deceleration 100 m
Speed (km/h)		20		34.2		34.2		13		20		20
Fuel consumption (litres/hr)		393.2		150		200		375		26.9		26.9


	Haul Truck Empty
Parameter	Acceleration 100 m		Flat (0%) Topo		Flat (0%) In-Pit/ Crusher/ Dump		Slope Up (10%)		Slope Down (-10%)		Deceleration 100 m
Speed (km/h)		25		45		45		26.9		25		25
Fuel consumption (litres/hr)		117.8		117.8		117.8		300		26.9		26.9

The calculated cycle times were based upon round-trip haulage profiles the haul truck speeds, and on load/spotpdump times determined for each material. A trend of each material type’s cycle time over the LOM is shown in Figure 16-14.

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LOGO

Figure 16-14: Cycle Time Trend over LOM

16.2.5.5

Equipment Annual Fleet Requirements

The primary mining fleet was selected based on the scale of this mining operation, optimization fleet size utilization and matching of equipment, efficiency and reliability. At the peak point in the mine life, primary equipment requirements will be as follows:

•

19 x 226-tonne diesel haul trucks;

•

2 x 28m³electric-hydraulic shovels;

•

1 x 28m³diesel-hydraulic shovel;

•

1 x 15m³front end loader; and

•

3 x 8¹/₂” DTH blast hole drills.

The haul truck fleet is shown in Figure 16-15 and follows the mined material trend over the LOM. The truck fleet takes also into consideration the units used during the construction of the tailings dam and during the PAG pile rehabilitation work.

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LOGO

Figure 16-15: LOM Haul Truck Fleet

To complement the primary mining equipment fleet, a list of auxiliary and support equipment was developed by BBA, and validated with Rainy River based on experience in similar open pit mining operations. The requirements for auxiliary support equipment were determined primarily based on the scale of the operation, the size and number of active waste rock piles and length of haul roads to be maintained.

Over the life of the operation, no mining equipment replacement is required. After Year 11, the final pit is depleted and most pieces of equipment are no longer needed. Hence, during the stockpile rehandling period, less equipment is required and their utilization is reduced significantly.

Table 16-13 shows the annual mine equipment fleet requirements over the open pit production stage to support the mining operation in each year.

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During the post-production stage, fewer equipment will be required to support activities related to stockpile re-handling, site rehabilitation and site maintenance. The necessary post-open pit equipment are: three (3) trucks (two (2) operating and one (1) back-up), one (1) large wheel loader, three (3) track dozers, one (1) motor-grader and one (1) compactor.

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Table 16-13: Annual Open Pit Mine Equipment Requirements


																									Stockpile Re-Handling
Mine Equipment	2014		2015		2016		2017		2018		2019		2020		2021		2022		2023		2024		2025		2026		2027		2028		2029		2030		2031
Haul Truck Fleet
Haul Truck (226-tonne)		6		6		9		15		19		19		19		19		19		19		19		7		7		3		3		3		3		3
Shovel Fleet
Hydraulic Shovel (28m³Diesel)		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1
Hydraulic Shovel (28m³Electric)						1		2		2		2		2		2		2		2		1		1		1
Drill Fleet
Blasthole Drill (8¹/₂” DTH)		1		1		2		3		3		3		3		3		3		3		3		1		1
Support Fleet
Wheel Loader (15m3)		1		1		1		1		1		1		1		1		1		1		1		1		1
Motor Grader (16”)		1		1		1		2		2		2		2		2		2		2		2		1		1		1		1		1		1		1
Excavator (520 HP)		1		1		1		1		1		1		1		1		1		1		1		1		1
Track-Dozer 580 HP		4		6		6		6		6		6		6		6		6		6		6		4		4		2		2		2		2		2
Track-Dozer 580 HP (for construction and rehabilitation)						1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1
Compactor						1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1
Auxiliary Fleet
Aggregate Plant		1		1		1		1		1		1		1		1		1		1		1		1		1
Boom Truck		1		1		2		2		2		2		2		2		2		2		2		1		1
Dewatering Pump (100 HP electric)		1		1		2		2		2		2		2		2		2		2		2		2		2
Fuel/Lube Truck (740FLT)		1		1		2		2		2		2		2		2		2		2		2		1		1		1		1		1		1		1
Hydraulic Crane, truck-mounted		1		1		1		1		1		1		1		1		1		1		1		1		1
Lighting Tower (4-post of 1000 W / diesel generator)		4		4		6		6		6		6		6		6		6		6		6		6		6		2		2		2		2		2
Mini Bus		1		1		2		2		2		2		2		2		2		2		2		1		1
Mobile Pump (150 HP)		2		2		4		4		4		4		4		4		4		4		4		2		2
Pickup Truck Crew Cab		6		12		12		12		12		12		12		12		12		12		12		6		6		3		3		3		3		3
DTH Drill (Reverse Circulation Sample, Pre-Split)		1		1		2		2		2		2		2		2		2		2		2		1		1
Scraper (used)		1		1		2		2		2		2		2		2		2		2		2		1		1
Service Truck		1		1		2		2		2		2		2		2		2		2		2		1		1		1		1		1		1		1
Stemming Loader		1		1		1		1		1		1		1		1		1		1		1		1		1
Tire Changer, truck-mounted		1		1		1		1		1		1		1		1		1		1		1		1		1
Tow Haul Truck (777D/F) (used)		1		1		1		1		1		1		1		1		1		1		1		1		1
Water Truck (740 – 30KL)						1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1
Water Truck (used 777D/F)		1		1		1		1		1		1		1		1		1		1		1
Small Fuel/Lube Truck (used)		1		1		1		1		1		1		1		1		1		1		1		1		1
Wheel Loader (used)		1		1		1		1		1		1		1		1		1		1		1		1		1

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16.2.6

Open Pit Mine Personnel Requirements

The personnel requirement for the open pit mine includes all of the hourly staff working in open pit operations that are required for the operation and maintenance of the equipment involved with or supporting mining activities, as well as the salaried engineering, geology and supervisory staff.

The number of hourly personnel reaches a peak of 248 in Year 2022 (Year 7 of production). A complete list of the hourly personnel requirements is listed in Table 16-14.

The maximum number of salaried employees is 51. The mine salaried staff requirements over the life of the mine are presented in Table 16-15.

The number of operators required for the major mining equipment (haul trucks, shovels, and drills) was determined according to the number of operating units and number of rotations during which the equipment is in operation. Most of the operators for the major mine equipment are based on a four (4) crew rotation. Hourly maintenance employee requirements were determined based on the number of equipment to maintain. The ratio of maintenance/operation for the hourly employees is approximately 0.61 during the normal years of operation. This ratio assumes that maintenance activities such as rebuilding components and machining will be performed off-site.

Post-operation activities (from Year 2026 to 2031), consisting of stockpile re-handling, environmental and site maintenance work, and personnel for operations and maintenance work will allow personnel requirements to be reduced accordingly. Consequently, after the open pit is depleted, the number of hourly and salaried staff will be reduced to a total of approximately 30.

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Table 16-14: Annual Hourly Personnel Requirements


	2014		2015		2016		2017		2018		2019		2020		2021		2022		2023		2024		2025		2026		2027		2028		2029		2030		2031
Operations
Shovel Operators		4		4		8		11		11		11		11		10		10		10		7		3		2		4		4		4		4		4
Loader Operators		4		4		4		4		4		4		4		4		4		2		2		2		2
Haul Truck Operators		18		21		33		55		63		68		70		72		78		73		66		25		20		8		8		8		8		8
Drill Operators		4		4		6		10		9		9		11		12		12		12		8		4		4
Dozer Operators		16		24		28		28		28		28		28		28		28		28		28		20		20		12		12		12		8		8
Grader Operators		4		4		4		8		8		8		8		8		8		8		8		4		4
Water Truck Operator/ Scraper		4		4		4		4		4		4		4		4		4		4		4		4		4
Other Auxiliary Equipment		4		6		6		6		6		6		6		6		6		6		6		4		4
General Labour		4		6		6		6		6		6		6		6		6		6		6		4		4
Janitor		2		2		2		2		2		2		2		2		2		2		2		2		2

Hourly Open Pit Operations Total		64		79		101		134		141		146		150		152		158		151		137		72		66		24		24		24		20		20

Field Maintenance
Field Gen Mechanics		4		4		9		13		13		13		13		13		13		13		11		4		4
Field Welder		2		2		4		7		7		7		7		7		7		7		5		2		2
Field Electrician		4		4		4		8		8		8		8		8		8		8		6		4		4
Shovel Mechanics		4		4		9		13		13		13		13		13		13		13		11		4		4
Shop Electrician		4		4		4		4		4		4		4		4		4		4		4		4		2
Shop Mechanic		10		12		14		14		18		20		22		24		24		22		20		8		8		2		2		2		2		2
Mechanic Helper		3		3		4		5		6		6		6		6		6		6		6		3		3		1		1		1		1		1
Welder-Machinist		2		4		4		6		6		6		6		6		6		6		6		4		4
Lube/Service Truck		2		5		5		5		5		5		5		5		5		5		5		2		2
Tool Crib Attendant		2		4		4		4		4		4		4		4		4		4		4		4		2

Hourly Mine Maintenance Total		38		47		61		79		84		86		88		90		90		88		78		39		35		3		3		3		3		3

Hourly Personnel Total		102		126		162		213		225		232		238		242		248		239		215		111		101		27		27		27		23		23

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Table 16-15: Salaried Open Pit Personnel Requirements


	2014		2015		2016		2017		2018		2019		2020		2021		2022		2023		2024		2025		2026		2027		2028		2029		2030		2031
Operations
Mine Superintendent						1		1		1		1		1		1		1		1		1
General Mine Foreman				2		2		2		2		2		2		2		2		2		2
Mine Shift Foreman		4		4		8		8		8		8		8		8		8		8		8		4		4
Drill & Blast Foreman				1		1		1		1		1		1		1		1		1		1		1		1
Blaster				1		2		2		2		2		2		2		2		2		2		1		1
Blaster Helper				1		2		2		2		2		2		2		2		2		2		1		1
Dispatcher				4		4		4		4		4		4		4		4		4		4		2		2
Training Foreman		1		1		1		1		1		1		1		1		1		1		1
Production / Mine Clerk				1		1		1		1		1		1		1		1		1		1
Secretary				1		1		1		1		1		1		1		1		1		1

Salaried Open Pit Operations Total		5		16		23		23		23		23		23		23		23		23		23		9		9		0		0		0		0		0

Maintenance
Maintenance Superintendent						1		1		1		1		1		1		1		1		1
Maintenance General Foreman				1		1		1		1		1		1		1		1		1		1
Maintenance Planner				2		2		2		2		2		2		2		2		2		2		1		1
Mechanical Engineer				1		1		1		1		1		1		1		1		1		1		1		1
Maintenance Foreman		1		2		4		4		4		4		4		4		4		4		4		2		2		1		1		1		1		1
Maintenance Trainer		1		1		1		1		1		1		1		1		1		1		1
Maintenance Clerk				1		1		1		1		1		1		1		1		1		1

Salaried Mine Maintenance Total		2		8		11		11		11		11		11		11		11		11		11		4		4		1		1		1		1		1

Engineering
Chief Engineer						1		1		1		1		1		1		1		1		1
Senior Mine Planning Engineer				1		1		1		1		1		1		1		1		1		1
Open Pit Engineer		1		1		1		1		1		1		1		1		1		1		1		1		1
Geotechnical Engineer		1		1		1		1		1		1		1		1		1		1		1		1		1
Blasting Engineer				1		1		1		1		1		1		1		1		1		1
Mining Engineering Technician		1		1		2		2		2		2		2		2		2		2		2		2		2
Mine Surveyor		1		2		2		2		2		2		2		2		2		2		2		1		1		1		1		1		1		1
Assistant Surveyor		1		2		2		2		2		2		2		2		2		2		2		1		1

Salaried Mine Engineering Total		5		9		11		11		11		11		11		11		11		11		11		6		6		1		1		1		1		1

Geology
Chief Geologist						1		1		1		1		1		1		1		1		1
Geologist				1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1
Grade Control Geologist		1		1		2		2		2		2		2		2		2		2		2		1		1
Geology Technician		1		1		2		2		2		2		2		2		2		2		2		1		1

Salaried Geology Total		2		3		6		6		6		6		6		6		6		6		6		3		3		1		1		1		1		1

Total Salaried Staff		14		36		51		51		51		51		51		51		51		51		51		22		22		3		3		3		3		3

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16.3

Underground Mining Methods

Golder Associates Ltd. (Golder) was responsible for the design of the underground mine for the Rainy River Gold Project. For the underground mine design, Golder used the SRK block model as well as the final open pit design provided by BBA. Golder was responsible for designing the underground mine, the infrastructure directly related to underground mining (surface fans, portals, backfill operations), and estimating the capital and operating cost of mining the mineralized material and transporting it to the surface stockpile. Surface facilities required for underground operations such as the shop, change-house, administrative offices, and water retention ponds were combined with facilities for the open pit operations and estimated by BBA.

The resources available to the RRU dip at approximately 45 degrees and vary from thicker zones (e.g., in the ODM area where mineralization can be more than 30 m thick) to thinner areas (e.g., in the 17 East area where mineralization is typically 5 m thick). Rock mechanics analyses indicate that the host rock is competent enough to allow open stopes over heights of 25 m. Due to the dip and varying thickness of the resources, it was decided that a combination of LHOS and CAF mining methods would be required.

Backfill for the Rainy River Underground operation will be a mixture of CRF in primary and longitudinal LHOS stopes, and un-cemented rock fill in the secondary LHOS and CAF stopes. This will reduce the quantity of cement required for the underground mine resulting in a lower operating cost, with a minimal impact on dilution.

Preliminary LHOS (Table 16-16) and CAF (Table 16-17) stope design parameters for each mining area were completed to help decide on the mining method to use in each area. The key differentials are vein dip and grade continuity. It was decided to limit the LHOS HW dip to a minimum of 60 degrees, as the blasted muck typically does not flow well along wall angles shallower than 55 degrees. Any areas that could not form an MSO envelope since the HW dip would be too shallow, or where the grade was not continuous enough, are mined by CAF. Another differential is mining cost. At this stage of the Study, it was assumed that the LHOS would be 20% cheaper than CAF to mine a tonne of ore. The total ounces (based on Au eq) available in each design were compared, and if the LHOS preliminary design provided 80% of the CAF ounces or more, then the

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LHOS mining method was chosen. Otherwise, the CAF method was chosen. A longitudinal section of the RRU is shown in Figure 16-16.

LOGO

Figure 16-16: Longitudinal Section of the Rainy River Resources Underground Mine

To limit the interaction between mining methods, each stope location was reviewed, and if access would make it impractical to mine, the method was changed. For example, if a small CAF area was surrounded by a LHOS area and it would be impractical to build the CAF attack ramps, then that area was switched to LHOS. The sub-level interval was set at 25 m for both methods to maintain consistency through the mine and to provide flexibility if changes to the method are required later. In all cases, ore will be moved to a re-muck bay on the level access and loaded into a truck for haulage to surface, so this was not a factor.

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Table 16-16: Preliminary Longhole Open Stoping Design Parameters


Item	Value		Unit
Hangingwall dip		60	Degrees
Sublevel interval		25	m
Cut-off grade		3.5	g/t Au_Eq
Primary stope width		10	m
Secondary stope width		20	m

Table 16-17: Preliminary Cut and Fill Design Parameters


Item	Value		Unit
Sublevel interval		25	m
Cut-off grade		3.5	g/t Au_Eq
Drift round size		5 x 5	m x m

16.3.1

Underground Geomechanical Design

The feasibility level site investigation work (AMEC 2013E) and underground mine design criteria for stope stability, ground support and backfill were performed by AMEC (2013A; 2013B; 2013C; 2013D; 2013G) and recommended to Golder Associates to support underground mining engineering.

During the 2012 drilling campaign (AMEC, 2013E), three main zones of the ODM/17 were intercepted: the West (BH12-UG-01), Central (BH12-UG-02) and East (BH12-UG-03) zones while the 433 UG North was delineated with the deeper borehole sections of BH12-OP-05 &-06.

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LOGO

Figure 16-17: View North West of the Underground Mine Geometry

Developed for the Map3D Numerical Stress Modelling, Indicating Zones and

Main Underground Boreholes (AMEC, 2013G).

In terms of rock mass quality, RQD’s were found to be excellent, ranging from 90% to 100% throughout all stoping domains and with respect to the Modified NGI Q-system, Q’ (after Barton et. al., 1974), average values of 23, 17, and 19, where obtained, characterizing the HW, OZ and FW domains of the largest west zone, respectively. The rock mass in all domains can be characterised as good (Barton et. al, 1974), however, there is a slight decrease in the quality in the central zone based on the present data.

Intact failure curves were developed based on laboratory testing of thirty (30) UCS, twenty seven (27) triaxial tests and twenty four (24) Brazilian tensile tests. The overall average UCS results, used for linear elastic numerical stress analysis using the Hoek-Brown brittle failure criteria (Martin et. al., 1999; Diederichs et. al., 2002; and Coulson, 2009) for the HW, OZ, FW and OZ+FW, and found to be 87, 125, 104 and 114 MPa respectively, indicating strong to very strong rocks.

Linear elastic stress analysis using the Hoek-Brown brittle failure criteria was used to review the sequencing, stress evolution around the development, and to estimate the levels of ground support required. Based on this and stope design using the Canadian Open Stope Stability Graph Method

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(Potvin, 1988; Hadjigeorgiou et. al., 1995), LHOS shallower than 500 m will not require cable bolt support to maintain hangingwall stability at shallow depths (Table 16-18). For LHOS deeper than 500 m and with transverse widths (ore thicknesses) greater than 15 m, cable bolt support is recommended in the back of secondary stopes, however, this accounts for only 7.5% of these deep stopes (AMEC, 2013G).

LOGO

Table 16-18: Modified Stability Graph for ODM West Zone Above 500 m Indicating Stability of Stope

Surfaces Based on Potential Design Limits and Actual Final Stope Dimensions

[Primaries 25 mH x 10 mL x 12 mW average dip 59 degrees] (AMEC, 2013G)

Ground support for the development at the RRU will consist of three (3) packages of increasing complexity (Table 16-19) depending on the location of the opening. Standard ground support will consist of resin rebar on a 1.2 m by 1.2 m pattern using #9 gauge wire mesh. The second and third packages account for development located in regions with increased damage due to mining induced stress. Most of the mine infrastructure will be supported with a standard ground support package, except for the deeper CAF stopes and some longhole stopes where increased levels of ground support have been assumed based on linear elastic modelling of the mine sequence. Overall, it is

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anticipated that approximately 5% of development will require shotcrete. Cable bolts are recommended in intersections with spans greater than 10 m.

Table 16-19: Ground Support Recommendations


Development Type	Maximum Profile	Code	Ground Support Levels
Development Type	Maximum Profile	Code	1	2	3
Standard Footwall Development Drives	4.5 mW x 4.5 mH	A	Back 2.4 m long-resin rebar on 1.2 x 1.2 m pattern with #9 gauge weldwire mesh Wall 1.5 m long-resin rebar on 1.2 x 1.2 m pattern	As per A1 except upgrade screen to #6 gauge for DL1¹ For DL2²add 60 mm of plain shotcrete	As per A2 plus based on DL3³ 5.4 m long Mn24 Swellex on 1.5 x 1.5m pattern in back and 2.4 m long Mn24 Swellex on a 1.5 x 1.5 m pattern walls

Intersections	> 10 m Span	B	As per A1 plus 6 m single Garford (GF) cables on a 2 x 2 m pattern	As per A2 plus 6 m single GF cables on a 2 x 2 m pattern	N/A

Ramp Passing Lane	10.5 x 20 m	C	As per A1 plus 3 central 8 m GF cables and 2 outside 6 m GF cables spaced 2 m on 2 m rings	N/A	N/A

LHOS Sill Drives	Deep Stopes Spans 20 mL x > 15 mW	D	As per A1 plus 6 m double GF cables on a nominal 2 x 2 m pattern fanned in back	As per D1 except upgrade screen to #6 gauge, replace wall bolts with 1.8 m long split sets (SS-39) for DL1¹ For DL2²add 60 mm of plain shotcrete	N/A

LHOS Draw Points	4.5 mW x 4.5 mH	E	As per A1 plus min 4 x rings, spaced 1 m apart of 3 m resin rebar OR 3 x rings of 3 x single GF cables spaced 2 m	N/A	N/A

CAF Sill Drives	5 to 8 mW x 5 mH	F	Back 3 m long 5/8” TBE rock bolts on a 1.2 x 1.2 m pattern with #9 gauge weldwire mesh Walls as per A1	As per F1 except upgrade screen to #6 gauge, replace wall bolts with 1.8 m long split sets (SS-39) for DL1¹ For DL2²add 60 mm of plain shotcrete	As per F2 plus based on DL3³ 5.4 m long Mn24 Swellex on 1.5 x 1.5 m pattern in back and 2.4 m long Mn24 Swellex on a 1.5 x 1.5 m pattern walls

CAF Sill Drives	> 9.5 mW x 5 mH	G	As per F1 plus 5.4 m long Mn24 Swellex on 1.5 x 1.5 m pattern in back	As per F2 plus 5.4 m long Mn24 Swellex on 1.5 x 1.5 m pattern in back	N/A

¹	DL1 = Linear elastic model induced stress = 0.33 s_c< s₁- s₃< 0.4 s_c(equiv. H-B brittle failure at m=0, s=0.11).

²	DL2 = Linear elastic model induced stress = 0.4 s_c< s₁- s₃< 0.5 s_c(equiv. H-B brittle failure at m=0, s=0.16).

³	DL3 = Linear elastic model induced stress = 0.5 s_c< s₁- s₃(equiv. H-B brittle failure at m=0, s=0.25).

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All primary stopes will be backfilled with cemented rockfill at an average binder dosage of 5%, while secondary stope will be filled with uncemented rock. Testing indicates that a 3% binder consisting of 10% cement (NPC) and 90% Slag has acceptable strength development at 28 days of 1 MPa (AMEC, 2013B).

Stope dilution has been based on the empirical estimation of wall slough after Pakalnis (2002) (Figure 16-18). Depending on the ore zones, stope HW dip and depth, 0.25 m to 0.5 m of ELOSs are estimated, for a sublevel spacing of 25 m, and 10 m and 20 m strike lengths for primary and secondary stopes respectively. CRF backfill dilution in secondary stopes from primary stope walls has been estimated at 0.3 m based on bench marking to similar operations (AMEC, 2013G).

LOGO

Figure 16-18: Empirical Estimation of Wall Slough (ELOS) for Varying HW dip Cases for the 4 Main

Underground Design Zones above 500 m depth in which LHOS was applied (AMEC, 2013G).

Based on the present mine plan, no crown or sill pillars will be required, since the stopes directly below the pit will be mined in advance, they are relatively narrow and will be backfilled on a primary secondary basis. The primary stopes will contain relatively stiff cemented rock fill, and as such are not anticipated to cause any instability to the open pit due to their limited extent.

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As some stress induced damage may be anticipated a 32-channel microseismic system is recommended for the deeper region CAF stopes and the region below the pit. Additionally, standard displacement monitoring and cable bolt monitoring instrumentation will be used.

16.3.2

Longhole Open Stoping

The LHOS areas were further divided into transverse and longitudinal mining zones depending on stope width: longitudinal in areas less than or equal to 8 m wide, and transverse for widths greater than 8 m.

16.3.2.1

Transverse Longhole Open Stoping

The typical sequence of a transverse LHOS mining block is shown in Figure 16-19, and is as follows:

•

Access drifts will be developed in the footwall at a distance of 20 m from the mineralized contact;

•

Cross-cuts will be developed from the FW access drift, and perpendicular to the strike of the ore body, to traverse the OZ and intersect the HW ore contact;

•

10 m wide primary stopes and 20 m wide secondary stopes in each stoping area will be planned on each level;

•

Primary stopes will be mined first. A typical LHOS stope with a drop raise requires four to five (4 to 5) drill and blast cycles to empty the stope;

•

Mucking will be manual until the brow opens, and then remote methods will be employed;

•

The maximum load-haul-dump (“LHD”) vehicle haul distance is kept to within about 250 m;

•

Once emptied, the primary stope will be filled with cemented rock fill with a 5% cement binder via the upper cross-cut for each individual stope. This fill becomes the mucking floor for the subsequent stope lift;

•

Secondary stopes will be mined adjacent to the backfilled primary stopes after sufficient fill cure time, which is typically more than 14 days;

•

Secondary stopes will be filled with unconsolidated backfill;

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•

Mining will progress from the bottom of a mining block upwards, and generally from the centre outwards; and

•

Primary stopes will lead the secondary stope mining by one level to ensure that there are no large stope back spans generated in the sequence.

LOGO

Figure 16-19: Schematic of the Typical Progression of Transverse Longhole Open Stoping

Typical stope development will consist of creating a slot raise towards the hangingwall side of the stope and drilling blast hole ring patterns parallel to the strike of the OZ and along the dip of the stope (i.e., dumped rings), in order to maximize ore recovery and keep dilution to a minimum. The drilling will take place in an overcut drift and the drop raise will allow the creation of a free face into which the blast holes will break.

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Production blast holes of 114 mm diameter will be drilled in a fan hole pattern using electro-hydraulic drill rigs. A typical drilling pattern for a transverse LHOS stope consists of a ring burden of 2.7 m and an average toe spacing of 3.3 m. Hole collars that are within 1 to 1.4 times the toe spacing will not be fully loaded, which will help to evenly distribute the explosive energy throughout the stope. A drop raise was favored for the RRU over a raise bore because of the lower equipment capital cost. The same drill rig can be used to drill the drop raise and the blast holes.

Blast holes can be charged with ammonium nitrate and fuel oil (AnFo) explosive in dry conditions or with bulk emulsion explosives when wet conditions are encountered. The powder factors for the primary and secondary stopes are calculated to be 1.5 and 1.2 kg/t, respectively, including the drop raise blasting. Approximately 72% of the Mineral Reserves at the RRU are within the category of transverse LHOS stope design.

16.3.2.2

Longitudinal Longhole Open Stoping

The typical sequence of a longitudinal LHOS mining block is shown in Figure 16-20 and is as follows:

•

Ore sills are developed:

•

to the far extents of the stoping area;

•

to follow the planned hangingwall stope contact to ensure grade control of the heading; and

•

at 4.5 m wide by 4.5 m high. If ore widths exceed 8 m, then transverse LHOS is employed, as previously described.

•

Mining commences from the far extent of the stoping areas and retreats back towards the access in stope lengths of up to 20 m depending on local geological conditions;

•

Mucking will occur on the sill below and the maximum LHD vehicle haul distance is kept to within about 300 m;

•

Pillars are not left between stopes unless an individual stope grade does not meet the cut-off grade criteria. In this case, the low-grade zone would simply be bypassed and left un-mined. Pillars have not been designed in the RRU; and

•

CRF is placed immediately in mined stopes via the upper sill for each individual stope. This fill becomes a stope wall for the next stope in the stoping area and the mucking floor for the subsequent stope lift.

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LOGO

Figure 16-20: Typical Longhole Open Stoping Drilling Layout (Ore Widths less than 8 m)

Longitudinal stopes will be drilled using the same equipment and hole size as the transverse LHOS stopes. Approximately 8% of the Mineral Reserves at RRU fall within this category of stope design.

There are four (4) stoping areas representing approximately 50,000 t that will require upholes. The drilling arrangements would be similar to the longitudinal LHOS, although the drill would be situated in the undercut with holes drilled up toward the upper sill.

16.3.3

Cut and Fill

At the RRU, 20% of the reserve is proposed to be mined by CAF methods. The typical mining sequence for a CAF mining block is shown in Figure 16-21 and Figure 16-22, and is as follows:

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•

Where possible, a drift is developed from the access ramp to the midpoint of a CAF mining horizon, which permits the mining of CAF lifts in two (2) opposing directions on any particular horizon to increase available working fronts and maximize production. Otherwise, the attack ramp will be connected to one end of the proposed stope lift;

•

The CAF access ramp is located between 100 m and 50 m from the mineralized zone to allow successive attack ramps to be developed up and down to a maximum 20% grade and follow the dip of the mineralization;

•

Each lift is 5 m high, and each sublevel contains five (5) lifts (two (2) down, one (1) at the same elevation, and two (2) up);

•

The maximum LHD vehicle haul distance is kept to within about 250 m;

•

Each lift is backfilled with unconsolidated backfill as tightly as possible to the drift back; and

•

Single drift CAF mining is done for the maximum mining width of 7 m. Ore widths up to 5 m will have standard support. Widths greater than 5 m will have increased support in the form of Swellex bolts as per ground support recommendations.

The excavation of each CAF lift will be done with conventional drill-and-blast tunnelling techniques, similar to the waste drift and ore sill development. Similar equipment will also be used.

LOGO

Figure 16-21: General Longitudinal Section of the Cut and Fill Mining Method

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LOGO

Figure 16-22: General Cut and Fill Ramp Arrangement

16.3.4

Backfill

Backfill is planned for the underground operations to promote stability and improve ore recovery. To limit cost, the RRU is designed with both cemented and un-cemented backfill. CRF was selected over paste fill due to the lower plant capital costs. The advantage of reducing the quantity of tailings deposited on surface was minimized as the portion of tailings originating from the underground mine at the RRGP is very small compared to the open pit.

Cemented rock fill will be used in the primary and longitudinal stopes, which account for approximately 40% of the underground longhole stopes. The CRF will be composed of 5% cement, 4% water, and 91% rock fill, and be placed by end-dumping with LHDs. Waste rock will come from either ore trucks that back haul from surface, or from development waste that will be stockpiled in the re-mucks. The cement will be mixed in a hopper attached to a spray bar positioned across the back of the access drift. The backfill LHD will collect a bucket of waste and pass under the spray bar to apply the cement. The cement-rock mixture will then be transported directly to the stope and dumped. The design criterion is that the CRF has to be self-supporting while the secondary stopes or longitudinal stopes are mined.

The secondary stopes will be backfilled with unconsolidated rock fill. This backfill cycle is similar to the CRF cycle, although there is no cement added. It was decided to use un-cemented rock fill in the secondary stopes as this backfill does not have to be self-supporting and it reduces the operating cost of the RRU.

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The backfill cycle for the CAF stopes will be similar to that of the secondary LHOS stopes. Waste rock will be stockpiled in the remuck on each level. Once a CAF lift is mined, the backfill LHD will move the un-cemented rock into the stope, tight-filling as much as possible. At regular intervals, the walls around the top of the drift will be painted. This will help the LHD operator on the next lift identify where the backfill ends and the ore begins. If this system results in too much dilution or ore loss, it may be necessary to use another method of backfill such as CRF, or to foam grout the top layer of the backfill prior to blasting the next level.

16.3.5

Underground Mine Design and Schedule

16.3.5.1

Mine Access and Egress

The RRU will be accessed by a main ramp that has a portal located to the east of the open pit as shown in Appendix H, close to the main crusher. It is designed to be 6.0 m wide by 5.5 m high, have an average grade of 15%, and passing lanes every 250 m to accommodate a high level of traffic and promote rapid development. The main ramp has a large cross-sectional area to accommodate large ventilation ducting required during construction (more information on the ventilation design is available in Section 16.3.7.1), and it is designed to be equipped with the main dewatering line from the underground.

The main ramp is also the sole means of access to the ODM, the lower 17 East Zones and the majority of the 433 Zone. The upper 17 East and a small pod in the upper 433 Zone will be accessed by portals in the pit.

The lower 17 East, 433, and ODM areas are accessed by secondary ramps that connect with the main ramp. These ramps are designed to be 4.5 m wide by 4.5 m high to accommodate the size of equipment planned for the RRU. Table 16-20 shows the design details of the mine access and egress.

Table 16-20: Design Details of the Mine Access Infrastructure


Item	Width (m)		Height (m)		Total Length (m)		Grade (%)
Main ramp		6.0		5.5		4,000		15
Passing lane		4.0		5.5		120		15
Secondary ramp		4.5		4.5		7,600		15

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Each mining zone is equipped with a secondary egress raise. These escape ways are designed to be 2.0 m in diameter and equipped with ladders and landings to allow workers to climb off the working level if the ramp is no longer useable during an emergency.

16.3.5.1.1

Refuge Stations

The RRU is equipped with three (3) portable refuge stations, one for each of the main mining areas. In addition, a permanent refuge station is planned near the warehouse/shop area. These refuge stations can act as lunchrooms during normal operating conditions.

16.3.5.2

Level Infrastructure

The level access consists of a main drift with an electrical substation, re-muck, ventilation access, and sump. The design details for each of these cut-outs are shown in Table 16-21. With the exception of the sump, the cut-outs are designed with a 2% grade to provide drainage. The sump is designed with a -18% grade to allow water flow. The re-mucks are designed to be high enough to allow loading or dumping of a 27 m³ truck and long enough to hold at least one truck load of material.

Table 16-21: Design Details for the Level Access Infrastructure


Item	Width (m)		Height (m)		Total Length (m)		Grade (%)
Electrical substation		5		4.5		800		2
Ventilation access		4		4		1,400		2
Remuck		5		6.5		400		2
Sump		5		3		100		-18

16.3.5.3

Footwall Drifts and Level Access

To limit the hauling distance of the LHDs, the footwall drifts have been designed to accommodate 27 m³ trucks and include a larger cross-sectional area for increased airflow. The footwall drifts are typically designed with a small ditch on the footwall side and a grade of between 0.5% and 2% towards the level sump for drainage. Once constructed, the majority of these drifts will be ventilated with flow-through ventilation, and ducting will not be required. Trucks will not be allowed

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in the footwall drifts requiring ducting. Table 16-22 shows a summary of the length and design details for the footwall drifts and level accesses.

Table 16-22: Design Details for the Footwall Drifts and Level Access

Item

Width (m)

Height (m)

Total Length (m)

Footwall drift and level access

4.5

10,500

16.3.5.4

Draw points and Cut and Fill Attack Ramps

The LHOS draw points and CAF attack ramps will be ventilated with auxiliary fans. They are designed to accommodate a 5.4 m³ LHD and a 1.0 m diameter duct. A summary of the design details is in Table 16-23.

Table 16-23: Design Details for Longhole Open Stoping Drawpoint

and Cut and Fill Attack Ramps


Item	Width (m)		Height (m)		Total Length (m)
LHOS draw points		4.5		5		9,000
CAF attack ramps		4.5		5		5,100

16.3.5.5

Material Handling

The majority of production mucking will be conducted using 5.4 m³ (8-yard) LHD vehicles. As most of the LHOS mucking will require remote operation, these LHDs will be configured with remote control packages. For the LHOS stopes, the LHDs will dig the ore at the draw points and tram it to a re-muck bay within the footwall access drift. Tramming distances for the LHOS will range from a minimum of 30 m up to a maximum of 310 m at the far end of the longitudinal stope sills, with an average of about 115 m. The same LHDs will load a fleet of three 27 m³ (60-tonne) trucks on each level, which will then haul the ore up the main ramp to the surface stockpile. In order to accommodate the LHD dumping height, a take-down-back (“TDB”) (a small area of the back that is slashed to make enough room to raise the bucket of the LHD and dump the load in the bed of the truck) will be required at each loading area.

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The majority of the waste rock will be mined from ramp and level development in advance of ore mining on a particular elevation. Development waste will be moved with 4 m³ (5-yard) LHDs and taken to a remuck bay located near the ramp intersections, where loading will occur, and hauled to the surface waste dump or the closest backfilling area using 15 m³ trucks. Similar truck loading areas with smaller TDBs will be required to move development waste.

16.3.6

Definition Drilling and Sampling

Definition drilling and sampling has been planned to properly delineate the longhole, and cut and fill stopes. In the longhole areas, it is envisioned that the first 5 m of each draw point will be excavated well in advance of production from that area. This cut-out will be used by the diamond drillers to properly define the stope boundaries. The CAF areas will be defined from drilling bays along the existing ramp system. In addition, each CAF round will be assayed to ensure a continuous grade.

16.3.7

Underground Mine Services

Underground mine services include infrastructure required to mine underground, such as ventilation, electrical and dewatering design, and water and compressed air requirements.

16.3.7.1

Ventilation

The ventilation design for the RRU was modelled using Ventsim Visual software. The airflow requirements were based on the Ontario Ministry of Labour standard, which states that at least 0.06 m³/s/kw (125 cfm) per bhp of power of the diesel equipment operating underground is required (Ontario Ministry of Labour 2013, Internet site). In accordance with these regulations, and the list of equipment presented in Table 16-24, it is estimated that 343 m³/s of airflow will be required at peak production. This estimate is considered adequate for a feasibility level design. However, it should be updated in the detailed engineering phase and continuously reviewed during operations.

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Table 16-24: Rainy River Ventilation Model – Peak Production Air Quantities


Equipment	Engine Size (kW)		Quantity		Unit Airflow (m³/s)		Utilization (%)
Jumbo drill		120		4		7.2		25
Longhole drill		120		1		7.2		25
Anfo loader		103		2		6.2		75
Emulsion loader		103		1		6.2		25
4 m³LHD		220		2		13.2		95
5 m³LHD		256		2		15.4		75
5 m³LHD - backfill		256		2		15.4		75
15 m³truck		293		2		17.6		75
27 m³truck		567		3		34.0		95
Bolter		115		5		6.9		25
Scissorlift		103		4		6.3		65
Boom truck		103		2		6.2		65
Grader		103		1		6.2		65
Jeep (including personnel carrier)		110		6		6.6		50
Shop/warehouse				1		27		100
Subtotal (m³/s)								298
Air Losses (5%)								15
Contingency (10%)								30
Total (m³/s)								343

Air is distributed throughout the mine depending on the quantity of production from each mining method and on the number of development headings. The quantity of air required for each activity was determined based on the typical equipment that would be located within the specified area.

16.3.7.1.1

Ventilation Design

At full production, the RRU will be ventilated with one (1) fresh air raise and two (2) exhaust raises, and the main ramp will be an exhaust airway. There will be multiple internal raises to distribute the fresh air and also act as escapeways. One temporary fresh air raise will be required during construction of the main ramp. Details of the raises can be found in Table 16-25.

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Table 16-25: Rainy River Ventilation Model – Raise Details


Raise	Final Air Type	Diameter (m)		Length (m)		Dip (degrees)
ODM – exhaust	Exhaust		4.5		570		65
Lower ODM	Fresh		2.5		170		65
Lower ODM escapeway	Fresh		2.0		110		65
ODM – fresh	Fresh		4.0		210		65
Upper ODM	Fresh		2.5		170		65
Upper ODM escapeway	Fresh		2.0		110		85
Fresh air	Fresh		4.5		440		65
17 East	Fresh		4.5		320		65
17 East – Escapeway	Fresh		2.0		270		65
433	Exhaust		2.5		310		70
Ramp – temporary	Fresh		2.5		280		89

Generally, the fresh air will flow into the mine through the fresh air raise, and the majority of it will be directed to the workplaces via the ODM fresh air raise, doors, wooden bulkheads, and small secondary fans. Exhaust air will be directed towards surface through the ODM exhaust raise. Sufficient fresh air for the anticipated truck traffic will be directed up the main ramp from the fresh air raise. When the 433 Zone is in production, fresh air will be directed into the 433 area via the main ramp. It will be exhausted out of a raise into the bottom of the pit, making the entire area flow-through ventilation. The 17 East area is connected to the bottom of the fresh air raise. Exhaust air from this area will exit through the 17 East and main ramps.

The Rainy River Gold Project ventilation system will be constructed in five (5) phases as the mine expands. A summary of the modelled airflow and pressures, and the estimated fan sizes and quantities is shown in Table 16-26. The longitudinal section of the Phase 3 design is shown in Figure 16-23.

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Table 16-26: Rainy River Ventilation Model – Fan Summary


Phase	Raise	Modelled airflow (m³/s)	Pressure (kPa)	Fan Size (kW)	Fan Quantity
1	Ramp – temporary	50	0.4	75	1
2	Fresh air	250	2.5	600	2
2	Upper ODM	60	1.0	110	1
2	Lower ODM	50	0.7	75	1
3	Fresh air	350	3.8	600	2
3	Upper ODM	10	0.05	75	1
3	Lower ODM	75	1.5	110	1
4	Fresh air	300	3.6	600	2
5	Fresh air	300	2.7	600	2

LOGO

Figure 16-23: Longitudinal Section of the Rainy River Underground Ventilation Layout where Red is

Exhaust Air and Blue is Fresh Air

The first phase will require a small 75 kW fan on the temporary ramp raise. The remaining four (4) phases will be served by the same fan and heater configuration on the fresh air raise. Two (2) 600 kW fans, both equipped with variable frequency drives, will be required to overcome the range of pressures estimated to be encountered throughout the life of the RRU. During production, the

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upper and lower ODM raises will require either a 75 kW or 110 kW fan, depending on which area has the majority of the activity. All the mine air will be heated by direct-fire propane heaters (32 MMBtu/h) attached to the fresh air raise fans.

For each phase, the quantity of air reaching the working face will be governed by bulkheads constructed in the ventilation access with variable openings, and by auxiliary ventilation. The majority of the longhole stoping in the ODM area will have flow-through ventilation with short ducting lengths. However, the CAF areas will require longer ducting lengths to bring the fresh air to the face.

Auxiliary ventilation will consist of a 75 kW fan attached to 1.0 m diameter ducting, which is standard in underground mines. The length of ducting will depend on the length of the heading, or the distance from flow-through ventilation to the workplace. The amount of air reaching the workplace will be greatly influenced by the quality of the ducting (the more rips in the ducting, the less air that will reach the face) and the installation of the ducting (ducting should be installed in messenger wire, and elbows used were appropriate to limit the friction losses). For longer ducting lengths, such as in the development to the ODM exhaust raise, zipper-type ducting should be used to reduce the losses caused by wire connections.

16.3.7.2

Mine Water Supply

It is estimated that the underground mining process will require 21.6 L/s of drilling water recycled from the mine. This will support all drilling and construction in the mine, and will be provided to the work area in 102 mm (4-inch) diameter HVC pipe installed during development of the footwall drifts and main ramp.

16.3.7.3

Dewatering

The dewatering system at the RRU will contain four (4) sumps. Each of the ODM, 433, and 17 East areas are designed with a sump to pump water to the main sump located near the warehouse/shop. From there, the water is pumped to the surface portal where it reports to the surface dewatering system. The size of each of the sumps is summarized in Table 16-27.

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Table 16-27: Summary of Sump Details for the Rainy River Underground


Sump	Pumping Height (m)		Flow (L/s)
ODM		235		36.6
17 East		185		20.6
433		40		36.6
Main		495		62.5

The dewatering estimate for each area includes the water consumed by the peak quantity of water-consuming equipment (jumbos, longhole drills, and bolters), estimates for dust suppression on the ramp, and on the muck piles, the estimated ground water inflow and a 25% contingency. AMEC estimated the average ground water inflow to be approximately 7.75 L/s, with model results ranging between 5.2 L/s and 18.9 L/s (AMEC 2012).

The mine plan includes having backfilled stopes near the pit bottom. It is conceivable that pit water may migrate through mining induced fractures in the pillar between the open pit and the underground. These potential inflows were not investigated nor included in the underground water modelling. The underground water inflow estimates and sump design should be updated in the next phase of work to consider inflow from the open pit into the stopes.

16.3.7.4

Compressed Air

It is estimated that the underground mining process will require 250 cfm of compressed air. This will support all drilling and construction in the mine. It will be provided to the work area from a compressor located near the portal in 152 mm (6-inch) diameter schedule 80 pipe installed during development of the drifts.

16.3.7.5

Underground Electrical

The electrical circuits for the underground mine and associated surface loads will be supplied from the main surface substation’s 27.6 kV GIS switchgear, as included in the BBA design. It is estimated that the RRU will have 6.6 MW in connected loads, resulting in a power demand of approximately 4 MW after efficiency, load and diversity factors are applied.

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BBA has included a 27.6 kV circuit to service the primary crusher. This circuit can be extended to the underground mine portal where it will supply a 27.6 kV breaker switchgear. This switchgear will supply a mine power centre (MPC) for the surface fans and air compressors as well as a MPC for the initial ramp development.

Once the main ramp reaches the 17 East ventilation drift and the associated ventilation shaft is prepared to surface, a new cable will be routed through the vent shaft to a disconnect switchgear located in the vent drift. This switch room will form the primary distribution centre for the mine. From this switchgear, a MPC will be fed that continues the main ramp development as well as MPCs for the 17 East ramp and production levels.

Once the main ramp reaches the 433 access drift, a new switch room will be prepared for a second disconnect switchgear. The main ramp development MPC will be fed from this switchgear.

Once the upper 17 East is ready for development, the breaker in the surface switchgear that previously fed the main ramp MPC will then feed a MPC for the upper 17 East.

16.3.7.5.1

Communications

For communications throughout the mine, a leaky feeder system is proposed. A reasonable number of readers and tags have been included for basic mine operations, however, additional units may be added, as required, for optimal mine monitoring. This leaky feeder system will provide two-way voice communications and limited data input for asset locations. Additional data beyond these requirements are not effectively supported by this type of system, therefore, if vehicle telemetry, video, and Voice-Over-Internet-Protocol (“VOIP”) communications are envisioned, then a modern fibre-based VOIP network is recommended.

Other systems have been included in the estimate, including instrumentation, control, SCADA, phone, and blasting systems.

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16.3.7.6

Underground Mine Maintenance and Warehousing

The RRU will be equipped with a small shop used for minor repairs (e.g., worn parts, tire, and hydraulic hose replacement/repair) and a warehouse. All major and preventive maintenance service will be completed in the main shop on surface; which will be shared with the open pit operators. The small shop and warehouse will be located near the access to the 433 mining area off the main ramp.

16.3.8

Underground Mine Schedule

The mine schedule was completed using Minesched, a plug-in to the Surpac software. The schedule is based on productivity estimates of the activities required to complete a particular task. For example, to complete a development round, the following activities must take place: drilling, blasting, mucking, scaling and bolting, and installing of services (i.e., ventilation tubing, and air and water pipes). To estimate the productivity of the development crew at the RRU, each activity was considered and the time taken to complete the activity, assuming no delays, was calculated. The continuous productivity was then factored to account for planned and unplanned equipment down time, workplace unavailability (due to issues such as ground control and ventilation), and shift breaks to estimate an average shift productivity for use in the development or production schedule.

16.3.8.1

Development

The development was scheduled in four (4) phases separated according to development rate. The first phase, between Q2 of 2016 and Q3 of 2017, involves a contractor excavating the main ramp. It is estimated that the contractor will be able to advance at 6.0 m per day (m/d) per heading. To improve the contractors ability to cycle equipment to the face (and traffic flow during production), passing lanes have been designed every 250 m, which will provide a convenient location to park idle equipment. Due to the passing lanes and ventilation cut-outs, the average advance rate used for this period is 6.5 m/d.

It is assumed that the contractor will have trained personnel and enough equipment to meet development targets. As the contractors work package includes the main ramp and small cut-outs along the main ramp, it is estimated that only one (1) jumbo crew will be needed. (One jumbo crew

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includes the jumbo, support equipment, LHD and enough trucks to efficiently move the material to surface).

The second phase assumes a multi-heading and multi-crew development approach between Q4 of 2017 and the end of 2019. This is the first phase of company-driven development. The productivity calculation previously discussed was completed for this phase, and it is estimated that the company development crews will reach three (3) rounds per day or the equivalent of 11 m/d. The calculations indicate that two (2) jumbo crews are needed to reach this development target.

Table 16-28 shows the estimated total development equipment required.

Table 16-28: Development Phase Equipment Requirements


Equipment	Quantity
Two-boom jumbo		2
LHD		2
Bolter		3
Truck		3
Scissor lift		2
Boom truck		2

Each heading has been limited to one (1) round per day or 3.7 m/d, with the highest priority heading being developed first. There is a ramp-up period of development productivity as RRR hires and trains employees. The contractor will still be on site for half of 2018 to make up any shortfalls in company development. There will be a peak in development in 2019 as the CAF equipment is purchased and initially used on capital development to increase the speed at which the initial production workplaces become available.

It is anticipated that during the third phase of development, from 2020 through 2024, the company development crew will be able to maintain three (3) rounds per day. The overall development rate will decrease from the previous phase as some equipment that was used for capital development is reassigned to CAF stoping.

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The fourth phase of mine development occurs between 2025 and 2028. This phase is characterized by declining development as the mine nears completion. The remaining areas include the operating development in the 433 and 17 East, which will be completed by the same crew responsible for the CAF stoping.

All vertical excavation not directly related to stoping will be done with a contract raisebore. It is assumed that the contractor will be able to provide a raisebore machine as required and achieve boring rates of 12 m/d.

The RRU is separated into three (3) main mining areas: the ODM, 433, and 17 East. The ODM Zone is targeted for initial production as it contains the highest grade material. However, proper ventilation infrastructure and accesses have to be installed prior to mining at full capacity from the underground. Therefore, the first development priority is to establish the fresh air raise, followed by the ODM exhaust raise and associated ventilation transfer drifts. Following this, the upper ODM will be developed to provide access to enough underground reserves to begin production while development crews excavate the ramp and the appropriate ventilation raises and escapeways to the bottom of the ODM.

Production mining in the 17 East area will begin after the CAF areas of the ODM. This requires the development crews to be shifted to this area one (1) year earlier to develop the ramp and raises. The 17 East area must be developed before the 433 to maintain constant CAF production.

Development crews will shift to the 433 area as soon as access, which is governed by the open pit mining schedule, is available. The 433 ramp will be excavated from the top (accessed through the bottom of the 433 OZ in the open pit) and the bottom (accessed from the main ramp). Open pit mining in the 433 area must be completed prior to installation of the underground infrastructure, and the 433 raise, which will be constructed in segments, must be fully established prior to production to provide sufficient airflow.

There are two (2) small satellite zones outside the main underground mine. The Upper 17 East is accessed off of the main pit at approximately 100 m elevation, and the Upper 433 area is

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accessed through the saddle in the open pit between the 433 and ODM OZs (approximately 155 m elevation). Both of these areas will be developed as they become available.

The longhole stope accesses are scheduled for development as they became available to allow proper hangingwall definition. The CAF attack level accesses will be excavated with the ramp, and are developed as needed.

Development of the RRU will produce 94.7 thousand tonnes of material grading above cut-off, with an average grade of 5.44 g/t Au. This material will be treated as ore and hauled to the mill. The RRU will also produce 844 thousand tonnes of incremental material (defined in Section 15.2.1.2) with an average grade of 1.44 g/t Au. This material will be either stockpiled underground for later milling, if space is available; used as backfill if required; or trucked out of the mine. Finally, it is estimated that the RRU will produce a total of 1.3 million tonnes of waste with an average grade of 0.08 g/t Au that will be used as backfill or hauled to surface depending on the availability of void space (see Section 16.3.8.3 for a discussion on the waste schedule).

16.3.8.2

Production

The production rates for the RRU were estimated for the different types of stopes using the productivity methodology discussed in Section 16.3.8 and are summarized in Table 16-30. The productivity of the CAF areas was estimated based on the position of the attack ramp. In the majority of cases, the attack ramp connects to the middle of the CAF lift (twin-access CAF lift). In some cases, due to the position of existing infrastructure, the attack ramp only connects to one end of the CAF lift. These areas are assigned a lower productivity than twin-access CAF lifts.

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Table 16-29: Stoping Productivities Used in the Mine Schedule


Area	Productivity (t/h)		Productivity (tpd)
Primary and longitudinal longhole stope		11		220
Secondary longhole stope		22		528
Double CAF lift		5.6		135
Single CAF lift		4.3		100

t/h = tonnes per hour; tpd = tonnes per day (22 working hours per day).

The RRU underground production schedule indicates approximately 750 tpd from the longhole areas, 250 tpd from the CAF areas, and a minor amount of ore from development. Overall, production from the RRU is scheduled to ramp up to approximately 1,000 tpd after five (5) years, produce approximately 1,000 tpd for seven (7) years and have one (1) year of reduced production at the end of the mine life.

Year 2021 shows a spike in production because of the ore from mine development. The RRU will require a significant amount of definition diamond drilling. To accommodate this, the longhole draw points are scheduled to be excavated as soon as they become available, which will result in an increase in ore-grade material for that period.

Overall, the RRU maintains an average production grade of above 5.0 g/t Au; the majority of the producing years have an average grade above 4.5 g/t. The underground mine has a peak grade of 5.68 g/t Au in 2022 as production comes from the main ODM Zone. Silver grade parallels the gold grade while mining on the edges of the ODM and the 433 Zones (at the beginning and end of the mine production) and increases significantly during production from the 17 East Zone between years 2024 and 2026.

16.3.8.3

Waste

The void space available underground was tracked and compared to the quantity of waste generated yearly. The difference represents the amount of backfill that will be required from surface. It should be noted that this is an idealized material balance, and it is difficult to schedule

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on a day-to-day basis due to the locations of waste generating development headings do not necessarily match with the stopes requiring backfill. Figure 16-24 shows the quantity of waste and void generated and the difference per period. Overall, it shows that the quantity of waste generated underground will be outpaced by the creation of voids once production starts. The RRU will require an average of 80,000 m³ (approximately four (4) truck loads per shift) of backfill to be hauled from surface for each year during production, assuming that all the waste generated during development on an annual basis can be used as backfill.

LOGO

Figure 16-24: Waste and Void Schedule from the Rainy River Underground (Golder 2013)

16.4

Combined Production Schedule

A combined production schedule for both open pit and underground operations is summarized in Table 16-30.

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Table 16-30: Open Pit and Underground Mine Production Schedule


	2014		2015		2016		2017		2018		2019		2020		2021		2022		2023		2024		2025		2026		2027		2028		2029		2030		2031		TOTAL
Open Pit Production
Mine to Mill (Mt)		—		—		2.87		7.66		7.66		7.52		7.42		7.25		7.33		7.31		7.30		7.31		0.49		—		—		—		—		—		70.12
Au (g/t)		0.00		0.00		1.19		1.31		1.35		1.72		1.58		1.36		1.12		1.10		1.15		1.39		1.39												1.34
Ag (g/t)		0.00		0.00		3.13		2.99		3.85		2.06		2.36		2.72		3.38		5.01		3.43		1.88		1.88												3.07

Mine COG (g/t)		0.6		0.6		0.6		0.6		0.6		0.6		0.5		0.5		0.45		0.4		0.45		0.3		0.3
Mine to Stockpile (Mt)		0.00		0.38		3.47		6.47		5.31		4.12		5.31		5.59		4.81		4.05		3.61		—		—		—		—		—		—		—		43.12
Au (g/t)		0.55		0.61		0.39		0.38		0.36		0.38		0.37		0.36		0.35		0.37		0.33																0.37
Ag (g/t)		1.74		2.03		2.12		1.94		2.23		1.19		1.78		2.04		2.15		2.39		1.80																1.97

Stockpile to Mill (Mt)		—		—		—		—		—		—		—		—		—		—		—		—		6.82		7.37		7.45		7.66		7.66		6.16		43.12
Au (g/t)																										0.37		0.36		0.36		0.32		0.30		0.52		0.37
Ag (gpt)																										1.98		2.17		1.40		1.92		2.10		2.28		1.97


Total Waste (Mt)		2.81		11.35		19.67		35.77		31.60		30.25		43.62		47.02		47.82		46.13		28.28		5.92		0.40		—		—		—		—		—		350.62

Total Overburden (Mt)		8.47		14.21		11.95		6.74		11.82		21.06		5.78		—		—		—		—		—		—		—		—		—		—		—		80.03


Total Moved (Mt)		11.28		25.94		37.96		56.64		56.39		62.94		62.13		59.87		59.96		57.49		39.19		13.23		7.70		7.37		7.45		7.66		7.66		6.16		587.02

Strip Ratio (waste:ore)						3.10		2.53		2.44		2.60		3.43		3.66		3.94		4.06		2.59		0.81		0.81												3.10

Underground Production
Cut and Fill Tonnage (Mt)														0.05		0.09		0.09		0.09		0.10		0.10		0.09		0.02		0.00								0.65
Au (g/t)														4.03		3.98		4.89		4.53		4.07		4.43		3.61		3.63		0.00								4.22
Ag (g/t)														2.23		1.86		1.90		14.26		38.49		41.70		30.62		25.80		0.00								20.56

Longhole Tonnage (Mt)												0.14		0.20		0.27		0.23		0.25		0.26		0.26		0.27		0.27		0.22								2.36
Au (g/t)												4.06		4.81		4.86		5.99		5.89		5.98		5.66		5.16		4.97		4.92								5.29
Ag (g/t)												4.15		4.80		3.30		2.96		2.30		2.00		2.02		2.06		3.05		4.90								3.04

Development Tonnage (Mt)										0.004		0.006		0.001		0.049		0.019		0.014		0.003		0.000		0.000		0.000		0.000								0.095
Au (g/t)										5.33		5.94		5.56		5.47		5.84		4.68		5.09		0.00		0.00		0.00		0.00								5.44
Ag (g/t)										5.69		6.81		2.65		2.59		3.43		2.21		1.42		0.00		0.00		0.00		0.00								3.06


Total Capital Horizontal Development (m)						1 200		3 000		5 550		4 800		4 1000		50		2 000		2 500		1 8000																3 995

Total Capital Vertical Development (m)						0		721		394.95		385		538		0		547		212		165																2 962


Total OP+UG (Mt) (Sent to Mill)		—		—		2.87		7.66		7.67		7.66		7.67		7.66		7.67		7.67		7.66		7.67		7.66		7.66		7.67		7.66		7.66		6.16		116.35

Au (g/t) OP + UG						1.19		1.31		1.36		1.76		1.68		1.54		1.32		1.31		1.35		1.58		0.64		0.53		0.49		0.32		0.30		0.52		1.08
Ag (g/t) OP + UG						3.13		2.99		3.85		2.11		2.43		2.73		3.35		5.03		3.85		2.41		2.31		2.27		1.50		1.92		2.10		2.28		2.76

¹	Open pit reserves have been estimated using a cut-off grade of 0.30 g/t equivalent Au and underground reserves have been estimated using a cut-off grade of 3.5 g/t equivalent Au.

²	Open pit reserves have been estimated using a dilution of 9.7% at 0.22 g/t Au and 1.31 g/t Ag and underground reserves have been estimated using a dilution of 9.7% at 1.56 g/t Au and 1.28 g/t Ag for the Longhole areas, and 9.0% at 4.16 g/t Au and 0.61 g/t Ag for the cut and fill areas.

³	Open pit reserves have been estimated using a mine recovery of 95% and underground reserves have been estimated using a mine recovery of 95%.

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17.

RECOVERY METHODS

17.1

Proposed Process Flowsheet

A flowsheet including crushing, grinding, gravity recovery, cyanide leach, carbon-in-pulp circuit, electrowinning and refining was developed based on metallurgical testwork conducted at SGS Lakefield, equipment suppliers and on BBA’s experience on similar projects. This flowsheet utilizes the results of the testwork completed to-date and is the basis for the plant design and mill operating cost developed in this Study.

The proposed process flowsheet for the Rainy River Project is shown in Figure 17-1.

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LOGO

Figure 17-1: Whole Rock Leach Process Schematic Diagram

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17.2

Process Design Criteria

The design and sizing of equipment was based on a total process plant throughput of 21,000 tpd, or 7.7 Mtpa. This includes 1,000 tpd from the underground mine starting in Year 3.

All equipment in the grinding circuit was designed based on the 80^th percentile of results from the comminution testwork program.

The general Process Design Criteria for the plant is shown in Table 17-1.

Table 17-1: General Process Design Criteria


Criterion	Unit		Value
Plant Availability		%		92
Throughput		tpa		7,665,000
		tpd		21,000
		t/h		951
Duration of Operation		years		16
Average Feed Grade (Open Pit and Underground, Blended, First 10 Years Operation)		g/t Au g/t Ag		1.46 3.19
Average Feed Grade (Open pit, excluding stockpile)		g/t Au g/t Ag		1.34 3.07
Average Feed Grade (Underground)		g/t Au g/t Ag		5.07 6.69
Average Feed Grade (Stockpile)		g/t Au g/t Ag		0.37 1.97
Average Feed Grade (Open pit and Underground, blended, life of mine)		g/t Au g/t Ag		1.08 2.76
Design Feed Grade		g/t Au g/t Ag		2.00 5.00
Primary Crushing
Number of Crushers				1
Crusher Type				Gyratory
Utilization		%		65
P₈₀		mm		163
Hourly Throughput		t/h		1,346
Grinding
Number of SAG Mills				1
SAG Mill T₈₀		µm		2,400
SAG Power Requirements		kWh/t		13.3
Pebble Crusher P₈₀		mm		13
Number of Ball Mills				1
Ball Mill P₈₀		µm		75

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Criterion	Unit		Value
Bond Ball Mill Index		kWh/t		15.0
Cyanide Leaching
Total Retention Time		hours		30
pH				10.5
Number of Tanks				8
Carbon-in-Pulp
Number of CIP Circuits				1
Carbon Tonnage per tank		t		20
Carbon Transfers per Day				0.5
Average Carbon Loading (Silver + Gold)		g/t		5,000
Number of Tanks				7

17.3

Process and Plant Facilities Description and Design Characteristics

The current design accounts for two (2) main mineral processing buildings:

•

Primary Crushing Building; and

•

Main Process Plant.

A general buildings site layout is presented in Table 17-2.

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LOGO

Figure 17-2: General Processing Area and Buildings Site Layout

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17.3.1

Primary Crushing

The primary crusher will be located on bedrock near the pit exit ramp outside the ultimate mine pit and outside the blasting perimeter. The pad will be approximately 200 m by 160 m to allow for mine truck circulation and turning. On the opposite side of the pad from the crusher will be a run of mine rock stockpile area.

Open pit mine trucks with a 226-tonne capacity and underground mine trucks with a 22 m³ capacity will dump rock into two (2) dump points, feeding a 1,372 x 1,905 mm (54” x 75”) gyratory crusher. The crusher will process 21,000 tpd, or 1,346 t/h at 65% utilization. The crusher will be powered by a 448 kW, 600 RPM squirrel cage induction motor.

A rock breaker will be used to break any large boulders and to manipulate rocks to avoid bridging the mouth of the crusher. Inside the crusher building, three (3) manual hoists will be used for general maintenance purposes. The auxiliary equipment located below the primary crusher is accessed through a hatch. A 20-tonne crane will be used for moving heavy equipment and pieces. A mobile crane will be used for moving the crusher spider and main shaft and for crusher liner replacement.

The primary crusher building is approximately 800 m from the process plant and houses the gyratory crusher and the tail end of the stockpile feed conveyor. The building foundation design and layout are based on using mechanically stabilized earth (“MSE”) to minimize capital costs. Also, a mechanically stabilized earth wall will be built so that the conveyor to the coarse ore storage remains aboveground. There are two (2) electrical rooms, with one located beside the crusher and the other located adjacent to the stockpile feed conveyor drive. The total depth of the building will be approximately 30.5 m below grade.

Crushed rock with an estimated maximum size of 350 mm (approx.14”) will be discharged to a surge pocket with a 330-tonne capacity. The surge pocket will serve as a buffer for the apron feeder during operation and will also be the access point under the gyratory crusher for maintenance.

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One (1) 2,134 mm wide variable speed apron feeder will reclaim crushed rock from the surge pocket and discharge onto the 1,372 mm wide crushed rock conveyor at a controlled rate. The apron feeder is sized to handle the full capacity of the gyratory crusher at 65% utilization. A dribble chute located under the apron feeder will be used to collect and discharge fines onto the crushed rock belt conveyor.

Sump pumps in the crusher and reclaim area will collect the underground water from outside the structure in addition to the water coming from operation.

17.3.2

Crushed Rock Handling and Storage

The crushed rock conveyor will discharge onto the stockpile feed conveyor (stacking conveyor) via a transfer tower.

The crushed rock stockpile pad is 75 m in diameter with a 20,000-tonne live capacity and an overall capacity of 60,000 tonnes. Under the stockpile, a reclaim tunnel is installed to recover the stored material. A prefabricated electrical and mechanical room is located adjacent to the reclaim tunnel, above ground at the exit point.

The crushed rock will be reclaimed by three (3) 2,134 mm wide apron feeders (three (3) running at a time with capability to run two (2) at a time), which will feed a single 1,372 mm wide SAG mill feed conveyor. A belt scale will be installed on a horizontal section of the conveyor belt to monitor the feed rate to the processing plant.

17.3.3

Processing Plant and Tailings Handling

The main processing building houses the grinding (SAG and ball mills), pebble crushing, gravity recovery, CIP, carbon stripping, electrowinning, refining and reagent preparation areas, as well as the tailings pumps, compressors and metallurgical laboratory. The pre-leach thickener, leach tanks, pre-detox thickener, lime slaking, offices and dry, and cyanide destruction areas are located outside of the processing building. Three (3) electrical rooms supply power to the plant.

The general plant layout is shown in Figure 17-3.

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LOGO

Figure 17-3: Process Plant General Arrangement Drawing (Including Primary Electrical Substation)

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17.3.4

Primary and Secondary Grinding

The 11.0 m x 6.1 m, dual-pinion 15,000 kW SAG mill will process an average 951 t/h of fresh feed with an F₈₀ of 163,000 µm. The SAG mill will be equipped with two (2) 7,500 kW motors and a variable speed drive system. Process water will be added to the mill to achieve a density of approximately 70% solids. Steel grinding media (5” or 125 mm diameter) will be used in the mill with a volumetric grinding media charge of approximately 13%. The discharge from the mill will be fed onto a single deck 3.6 m x 7.3 m scalping screen to size the ball mill circuit feed. The oversize from the sizing screen will be fed to the pebble recycle conveyor located at the discharge of the screen, which then feeds another recycle conveyor to a 448 kW pebble crusher. The sizing screen oversize tonnage will be approximately 25% of the fresh feed. The fresh feed from the stockpile will be combined with the crushed pebble recycled from the pebble crusher to feed the SAG mill.

The scalping screen undersize, with a T₈₀ of 2,400 µm discharges into the cyclone feed pump box.

The slurry from the cyclone feed pump box will be pumped to a cyclone cluster in closed-circuit with a 7.9 m x 12.3 m, 15,000 kW ball mill. The overflow from the cyclone cluster will feed the pre-leach thickener via two (2) 20 m² linear trash screens placed in parallel to remove any unwanted material and will have a P₈₀ of 75 µm. The cyclone underflow will be fed to the ball mill and the overall circulating load of the ball mill closed circuit is estimated to be 300%. A fraction of the Ball mill discharge will be diverted to the gravity recovery circuit.

The ball mill will be equipped with two (2) 7,500 kW motors (dual-pinion) and a variable speed drive system. The variable speed on the ball mill will permit closer control on the product size and will also permit a soft start for frozen charge control. The drive size was chosen to match the SAG drives, as this was the most economical sizing option.

Sufficient space shall be provided behind and around the ball and SAG mill to use a mill liner handler for effective mill maintenance. The mill liner handlers will be the hydraulic driven type. There will be a hydraulic inching drive provided for both the SAG and ball mills to rotate the mills to the desired position during liner maintenance.

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The grinding building is a one-bay building covering all the grinding facilities, containing the SAG and ball mill, along with the sizing screen, pebble crusher, cyclone cluster and gravity recovery equipment. This area is serviced by one (1) overhead crane with a 60/20-tonne capacity to lift the heaviest pieces of the mills. The SAG mill electrical room is located on the ground floor under the SAG mill feed area.

Grinding media is stored in three (3) ball pits located along the exterior wall beside the SAG mill. Balls are retrieved from the ball pit by means of the 60/20-tonne capacity crane through a ball flow gate system and ball bucket. The grinding media is fed into the SAG and ball mill through ball addition chutes.

17.3.5

Gravity Circuit

The feed to the gravity recovery circuit is anticipated to handle approximately 600 t/h. The gravity feed will be a portion of the ball mill discharge which is diverted from the ball mill discharge chute to a dedicated gravity circuit feed pump box and pumps. Two (2) 1,800 x 4,900 mm screens will scalp off any coarse material prior to the gravity concentrators. The undersized material from the screens will feed two (2) 56 kW gravity concentrators, placed in parallel. The gold concentrate from the gravity concentrators will feed an intensive cyanidation vessel. The pregnant leach solution from the cyanidation vessel will be pumped to a dedicated electrowinning cell via a pregnant solution tank located in the gold room. The gravity concentrator tails will be returned to the ball mill pump box. The intensive cyanidation tailings will be sent to the leach tanks.

17.3.6

Cyanide Leaching Circuit

A 44 m diameter pre-leach, above-ground thickener positioned outside of the concentrator building will increase the density of the ball mill cyclone overflow to approximately 62% solids. Lime (CaO) will be added to the thickener feed box to raise the pH of the slurry to around 10.5 to improve the solids settling rate. The underflow from the thickener will be diluted to 50% solids with cyanide-bearing process water prior to being fed to the leach tanks by slurry pumps located under the thickener. The thickener overflow will flow into the 17 m diameter non-cyanide bearing process water tank located between the thickener and the process plant. The process water pumps will be located inside the plant. The thickener drive mechanism will be enclosed in a shelter to prevent

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snow, ice and other weather hazards from damaging the equipment and to facilitate operation and maintenance during winter.

Gold leaching will be performed using sodium cyanide (“NaCN”). Lime slurry will be added as required to maintain the pH of the solution around 10.5-11. Eight (8) 18 m diameter agitated leach tanks will be used in a series arrangement, allowing for a 30-hour retention time. The heights of the tanks will range from 22.9 m to 18.0 m, with the slurry flowing from the tallest tank to the shortest. Oxygen will be used in the leach tanks to improve reaction efficiency and pacify sulphides. While no improvements in kinetic were noted using oxygen over air in testwork (refer to Section 13.8.4), a trade-off study indicated that using oxygen was more economically favourable than compressed air. The leach tanks will be installed on concrete rings resting on rock. A concrete slab and concrete wall poured on a granular backfill will serve as secondary containment around the tanks. Any seepage through the steel bottom of the tanks and concrete slab will be contained by a high-density polyethylene (“HDPE”) membrane installed in the backfill. Tanks and agitator drives will be serviced by a mobile crane.

17.3.7

Carbon-in-Pulp Circuit, Carbon Stripping and Reactivation

A carousel style CIP circuit was selected. Seven (7) 330 m³ tanks will be required, with each tank containing 20 tonnes of carbon. The retention time for each tank will be approximately 15 minutes, with approximately 112 minutes total retention for the circuit.

With the carousel system, it is estimated that the maximum carbon loading will be approximately 10,000 grams of gold and silver per tonne of carbon (g/t), with average loadings of approximately 5,000 g/t. One (1) tank from the CIP circuit will be emptied and transferred to the stripping circuit every two (2) days. The contents of the tank will be pumped to the loaded carbon recovery screen located on top of the carbon stripping circuit. The carbon retained on this screen will feed the stripping circuit while the slurry and wash water undersize will be recycled to the CIP circuit.

The strip solution will exit the barren solution tank and will be heated via heat exchangers and immersion heaters. This solution, containing sodium cyanide and caustic soda, will capture gold from the carbon located in the stripping vessel and will be cooled down via the same heat exchanger arrangement prior to electrowinning. The electrowinning circuit will produce a gold and

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silver sludge, which will be dried and smelted into doré bars as a final product. The now barren solution will then return to the barren solution tank.

The stripped carbon discharged from the stripping circuit will be screened: the oversize will be reactivated in a kiln and recycled to the CIP circuit, while the fines will be recovered and sold to a refinery for gold credits.

Fresh carbon will be added into a carbon attrition tank via a hopper. Carbon transport water will be added to the tank and the tank’s agitator will condition the carbon. The carbon slurry will be pumped to a fresh carbon sizing screen. The oversize from the screen will be discharged into the quench tank while the undersize will be discharged into the fine carbon collection tank.

17.3.8

Tailings Management and Cyanide Destruction

The CIP tailings will be pumped to a 44 m diameter pre-detox thickener via a 32 m² safety screen, allowing recovery of carbon that may have passed through a broken carbon retention screen. The overflow from the thickener will be stored in the cyanide bearing process water tank and used as cyanide bearing process water predominately in the dilution of the pre-leach thickener underflow. The pre-detox thickener underflow, discharged at 60% solids will be diluted to 50% solids using non-cyanide bearing process water. The diluted underflow will be pumped to the cyanide destruction circuit. The cyanide destruction circuit will use a conventional liquid SO₂/air process to lower the weak-acid dissociable cyanide (“CN_WAD”) and total cyanide (“CN_T”) levels to acceptable levels for discharge into the tailings pond. Sulphur dioxide (SO₂) will be added along with oxygen (in the form of compressed air) to dissociate the cyanide, along with dissolved copper acting as a catalyst. Hydrated copper sulphate (“CuSO₄•5H₂O”) will be added to raise copper levels in the solution and act as a catalyst. Lime will be used to neutralize any H₂SO₄ produced in the cyanide destruction reaction. The cyanide destruction circuit will require one (1) 13.5 m x 16 m tank in order to provide 90 minutes of retention time. The cyanide destruction tank will be located outside the process building, installed on a concrete ring resting on rock. A concrete slab and concrete wall poured on a granular backfill will serve as secondary containment around the tanks. Any seepage through the steel bottom of the tanks and concrete slab will be contained by a HDPE membrane installed in the backfill. The product from the cyanide destruction circuit will be pumped to the

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tailings pond. Water from the tailings pond will be reclaimed and pumped into the process as non-cyanide bearing water.

A single pipeline will be used for tailings pumping. The pipeline will be approximately 6.5 km in length and made of HDPE. There will be a set of standby tailings pumps that can be used when required. Reclaim water will be returned via a HDPE pipeline with the use of a floating reclamation barge.

In case of an extended power outage, the contents of the tailing pipe lines will be flushed with reclaim water and then drained by gravity into two (2) outside emergency retention ponds to prevent settling of the slurry and pipeline blockage.

17.3.9

Refining Area and Gold Room

The refining area and gold room will be a secure area with two (2) separate entrances. The first entrance is a personnel entrance to the gold room office. The personnel entrance consists of a reinforced security door which leads into a staging room. The staging room leads into the gold room through another reinforced security door, which can only be opened when the first security door is locked. The second entrance is for the armoured truck to enter the gold room. An external safety fence is erected around the reinforced garage door and is setup in the same way as the personnel entrance, where the security fence gate must be locked before the reinforced garage door is opened.

This area will contain the electrowinning circuit, the induction furnace to produce the doré bars, a vault, and the gold room office. It is anticipated that during full production, an average of 12 doré bars will be poured per week.

17.3.10

Reagent Areas

All process reagents will be located in a separately contained area within the process plant building to prevent contamination of the plant in case of a spill. The reagent area will also include the exterior lime storage bin.

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17.3.11

Control Room and Maintenance Shop

The control room will be located in the grinding area on an elevated floor, overlooking the SAG and ball mill circuits. The plant maintenance shops will be located on the ground floor in the grinding area. This area will be serviced by a 10-tonne capacity overhead crane.

17.3.12

Offices and Change House

The process plant staff offices will be located outside and adjacent to the process plant in a prefabricated building. This building will house the conference room, documentation room, computer server room, lunch room and washrooms. Change rooms for men and women will be located in the prefabricated building.

17.3.13

Metallurgical Laboratory

A metallurgical laboratory will be included in the process plant. The laboratory will be located on the eastern side of the plant, adjacent to the reagent storage area.

17.4

Energy, Water and Consumable Requirements

17.4.1

Energy

The power demand of the process plant will be approximately 41.76 MW resulting in an energy consumption of 43.9 kWh/t milled for 21,000 tpd. The grinding circuits represent approximately 60-65% of the total operating power of the plant. The processing plant power demand is shown in Table 17-2.

Table 17-2: Process Plant Power Demand by Area


Area	Power Demand (MW)¹
Crushing (Gyratory and Pebble Crusher)		0.50
Grinding (SAG and Ball Mill)		25.72
Processing (Peripherals and Back end)		11.00
Network Loss (2%)		0.74
Power Demand Subtotal		37.97
Security Factor (10%)		3.80
Process Plant Power Demand²		41.76

¹	The power demand was calculated using various efficiency, load and diversity factors.

²	Tailings barge is not included in total process plant power demand. Refer to Section 18 for the whole site power demand.

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17.4.2

Water

A water balance has been developed and is shown in Figure 17-4. The addition of cyanide into the circuit will initially be done after the pre-leach thickener; however, the design allows for cyanide to be added into the grinding circuit in the future, if required. The overflow from the pre-leach thickener will be combined with reclaim water from the TMA and MRP to make up the non-cyanide bearing process water. The MRP serves as a catchment basin for the mine waste rock pile.

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LOGO

Figure 17-4: Process Plant and Tailings Ponds Water Balance

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It can be seen that process water will be obtained from both the TMA (reclaim water) and from the MRP. Approximately 40% of the make-up process water will come from the TMA. The reclaim water will be used as gland seal water, heat exchanger water for cooling of major pumps and make-up to the process water tank. The remaining process make-up water will be taken from the MRP. Overflow from the Stockpile Pond (“SP”) will also be pumped and collected in the MRP. The total process water make-up has been estimated at approximately 875 m³/h.

Fresh water will be obtained from the West Creek Pond (“WCP”) and will be used for reagent preparation, surface utilities and fresh water plant requirements. The total fresh water requirement is estimated to be approximately 75 m³/h.

17.4.3

Consumables

Reagents will be required for various areas of the plant such as: cyanide leaching, CIP, stripping and refining, thickening and cyanide destruction.

Lime

Quick lime will be delivered in trucks with 30-tonne capacity into a 183-tonne storage silo, sufficient for 10-day capacity. Volumetric screw feeders at the bottom of the silo will convey the quick lime to two (2) lime slaking trains (one (1) operating, one (1) stand-by), where water is added to the quicklime to form a hydrated lime slurry. The 130 m³ holding tank will have a capacity of 1.5 days. The hydrated lime slurry will have a solution strength of 25% w/w (mass fraction weight/weight).

The lime slurry will be fed by two (2) distribution loops to the grinding circuit, pre-leach thickener, leach tanks and cyanide destruction circuit. A by-pass line to the leach tanks will be provided in the event of a pH variation.

The total autonomy of the system will be approximately 12 days.

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Cyanide

Sodium cyanide will be delivered as solid briquettes in 20-tonne capacity ISO containers. The briquettes will be mixed with water and caustic soda to form a 23% w/w solution in a 100 m³ mixing tank. Two (2) recirculation pumps will be provided to recirculate the solution with the ISO container to ensure proper dissolution of the sodium cyanide. A 150 m³ holding tank will also be provided. Two (2) variable speed pumps will deliver the cyanide solution into the leach circuit, stripping circuit and intensive cyanidation unit using flowmeters as control elements.

The total autonomy of the system will be approximately seven (7) days. One (1) extra ISO container is left on-site, allowing an extra three (3) days autonomy.

Oxygen

A turnkey oxygen plant will be installed and operated on site by a third party beginning in Year 3 (2018). During the first two (2) years of operation, oxygen will be supplied as bulk liquid and vaporized on site. This will be done due to the lower tonnages expected in Year one (1) prior reaching steady state and to gather a better understanding of the oxygen requirements prior to installing an on-site oxygen generator (“VPSA”).

The vacuum pressure swing adsorption VPSA plant will be capable of supplying approximately 750 m³/h of oxygen. A liquid oxygen backup system, as used in Years 1-2, will also be integrated into the package. This arrangement will ensure a constant gaseous oxygen supply to the leach tanks. All oxygen piping will be stainless steel.

Caustic Soda

Caustic soda will be delivered in liquid form (50% w/w) in tanker trucks of 30 tonnes. A 50 m³ mixing tank will be provided to dilute the caustic soda solution to 30% w/w. The holding tank will have a volume of 70 m³. The caustic soda will be pumped to the cyanide mixing tank, the stripping circuit and the intensive cyanidation unit.

The total autonomy of the system will be approximately 30 days.

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Anti-Scalant

Anti-scalant will be used in various areas to minimize the scale build-up. Each area will have its own anti-scalant metering pump. Anti-scalant will be delivered in 30-tonne tankers and stored inside in a 43 m³ reservoir.

The total autonomy of the system will be approximately 140 days.

Nitric Acid

Nitric acid will be delivered at a concentration of 67% w/w by a 30-tonne tanker truck and transferred to the 37 m³ nitric acid holding tank. The nitric acid will be diluted and used in the stripping circuit for acid washing.

The total autonomy of the system will be 30 days.

Carbon

Natural coconut shell-type activated carbon (typical dimensions 6 mesh x 12 mesh) will be used in the adsorption circuit. The total estimated consumption will be approximately 30 g of carbon per tonne milled, based on operation standards and the utilization of the carbon-in-pulp pump cell circuit minimizing carbon losses as fines. Monthly consumption will be approximately 20 tonnes of carbon. Carbon will be delivered in super bags and stored outdoors.

Sulphur Dioxide

Sulphur dioxide (“SO₂”) will be delivered in liquid form by a tanker truck of approximately 38 m³ and stored in a 64 m³ pressurized horizontal holding vessel. The package will be complete with a padding system assembly including air compressors, dryers and compressed air receiver. The padding system will ensure storage of sulphur dioxide in its liquid form by pressurizing the horizontal vessel. The padding system will also allow liquid sulphur dioxide delivery to the cyanide destruction tank. Package will be complete with all required instrumentation for metered reagent delivery. The selected arrangement ensures that no compressed air lines connected to the SO₂ system enter the process plant.

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The total autonomy of the system will be 10 days.

Copper Sulphate

Copper sulphate (“CuSO₄•5H₂O”) will be delivered in 1,000 kg super bags. The bags will be mixed with fresh water and dissolved to 10% w/w. The 36 m³ mixing tank will be sized for 1.5 days of production. The solution will be transferred from the mixing tank to a 54 m³ holding tank by a transfer pump, from where it will be pumped to the cyanide destruction tanks via metering pumps.

The total autonomy of the system will be approximately four (4) days. The on-site storage of solid copper sulphate in super bags will ensure an autonomy of approximately 30 days.

Flocculant

Flocculant will be delivered to the plant in 750 kg super bags. The two (2) thickeners will require approximately two (2) bags per day. The flocculant bags will be stored in an outdoor container. A small supply will be kept in the mixing area. The bags are lifted over a hopper which feeds a wetting device to form 0.75% w/w slurry. The slurry will then report to an agitated mixing tank prior to being transferred to a flocculant holding tank by progressive cavity pumps. The mixing tank was sized to have a 4-hour capacity.

From the holding tank, the flocculant will be metered independently to both thickener feed wells by progressive cavity pumps. As the polymer is pumped by the metering pumps, it will pass through a flocculant dilution board (static mixers), where it will be diluted further to 0.05% w/w. The holding tank will have a 24-hour capacity.

The total autonomy of the system will be 28 hours. The on-site storage of flocculant in super bags will ensure an autonomy of approximately 30 days.

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Reagent Consumption

A breakdown of the estimated reagent consumptions per tonne milled and total annual consumption is presented in Table 17-3.

Table 17-3: Process Plant Reagent Consumption


Reagent	Formula	Physical State	Estimated Consumption
Reagent	Formula	Physical State	g/t milled		tpa
Lime	CaO	Solid		820		6,260
Sodium Cyanide	NaCN	Solid		280		2,150
Oxygen	O₂	Liquid¹		650		5,000
Caustic Soda	NaOH	50% w/w Solution		120		910
Sulphur Dioxide	SO₂	100% w/w Liquid		390		3,000
Copper Sulphate	CuSO₄•5H ₂O	Hydrated Crystals		70		520
Nitric Acid	HNO₃	67% w/w Solution		60		430
Carbon	—	Solid		30		260
Anti-Scalant	—	Solution		15		100
Refining Fluxes	—	Solid		1		8
Flocculant	—	Solid		70		550

¹	Oxygen to be received as liquid during first two years of operation with on-site oxygen generation installed in Year 3 of operation.

The reagent consumptions were based on project specific testwork, supplier recommendations and operating practice in existing plants.

Grinding Media Consumption

Other consumables include the grinding media for the SAG mill and ball mill. The type, size and estimated consumption of grinding media by piece of equipment are shown in Table 17-4.

Table 17-4: Grinding Media Consumptions by Mill Type

Equipment

Type

Size
(mm)

Estimated Consumption

g/kWh

g/t

tpa

SAG Mill

Forged Steel

125

650

5,000

Ball Mill

Forged Steel

50-75

750

5,750

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18.

PROJECT INFRASTRUCTURE

Layout drawings of the general infrastructure, including the process plant, offices, administration building, main electrical substation, truck shop, truck wash facilities, coarse ore storage and crusher building are provided in Appendix H.

18.1

General Site Works

General site preparation will consist of clearing, grubbing, topsoil removal and surface leveling throughout the construction areas. Clearing, grubbing and topsoil removal needs were estimated with aerial photographs showing tree and ground cover. Topsoil removal and grubbing are considered to be removed at the same time for material take-off purposes. Clearing is done in and around all construction areas to provide easy access. Topsoil is removed to provide a stable sub-base for platforms and to provide slope stability below the overburden and waste rock stockpiles.

Site drainage will be achieved with the excavation of drainage ditches along the building platforms, roads, culverts and sedimentation ponds.

Underground sanitary sewers, underground fire protection and potable water pipes, as well as sewage and potable water treatment plants will be constructed according to local requirements. Potable water will be distributed to the process plant area and the mine garage. These areas will also have sanitary sewers and fire protection. All areas will have a granular access platform surrounding the facilities.

Engineered fills on clay foundations must have at least 3H:1 V side slopes to avoid over-stressing the clay. Excavations deeper than 2 m have slopes no steeper than 3H:1 V for stability and safe working conditions. A frost depth of 2.7 m is also to be considered for building foundations or underground piping that are not sitting on rock.

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18.1.1

Primary Site and Access Roads

Site access roads to the TMA and to the explosives plant are already existing roads. The road widths will be enlarged to allow space for tailings and reclaim water pipes, as well as light traffic (emulsion tankers and pickup trucks). Existing roads will be resurfaced with crushed stone.

Plant site roads connect the process plant area to the coarse ore storage and the crusher area. The roads represent a total length of approximately 900 m.

TBT was commissioned by Rainy River to undertake a study for construction of an east access road that will serve as the main access from Kings Highway 71 to the proposed mine development area and provide maintenance access alongside the tailings pipes.

The TBT study also determined the optimal route for the proposed realignment of a segment of Provincial Highway 600, currently passing through the development area. Refer to Appendix H for a map of local roads surrounding the Project site and the proposed Highway 600 reroute.

Based on the findings of the TBT study, the preferred alignment for rerouting Highway 600 around the proposed development area optimizes the use of existing road easements and is the preference of both the Township of Chapple and Rainy River. Access to Marr Road will be provided along the bypassed section of old Highway 600 from Marr Road westerly to Pine River Road (new Highway 600).

Further verification of ground terrain and soil conditions are required in subsequent stages of project development to confirm detailed quantities and cost estimates.

18.1.2

Mine Haul Roads

Mine haul roads will be built to connect the open pit to the overburden and waste rock stockpiles. These haul roads will also connect the pit to the crusher pad, mine facilities (truck shop and truck wash) and tailings dike. The total length for the mine haul roads outside the pit limit is approximately 5,400 m. These roads will be built at the start of the Project and will remain in use for the duration of the mine life.

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18.2

Geotechnical

The geotechnical site investigations for the purposes of slope and dam design included 16 boreholes for the tailings dams, five (5) boreholes for the overburden stockpile and five (5) boreholes at the mine rock stockpile. An additional 26 boreholes were put down to allow for the design of the open pit overburden slopes, to obtain samples for characterization of the mine waste overburden for use as dam construction material, and to determine the hydrogeological conditions between the pit and the Pinewood River.

The geotechnical investigations for the process plant, carried out under BBA’s specifications and requirements, included a comprehensive drilling and bedrock depth probing program consisting of two drilling campaigns comprising 34 geotechnical boreholes, 38 test pits and 73 dynamic cone penetration tests. An iterative process between AMEC and BBA was followed in order to place the process plant facilities on bedrock.

AMEC (2013a) provides further details on the geotechnical and hydrogeological investigations. The scale of drilling and bedrock depth probing campaign is considered adequate for the feasibility level design of the facilities.

The foundation design recommendations for the plant facilities provided in AMEC (2012b) were incorporated into the design of the facilities by BBA.

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18.3

Mine Services Facilities

The mine services facilities for both the open pit and underground operations include the mine garage and the truck wash facility. The facility dimensions are based on a typical 793F haul truck. The overall vehicle dimensions and recommended repair bay specifications are summarized in the Table 18-1 and Table 18-2.

Table 18-1: Mining Vehicle Dimensions


Dimension	793F Haul Truck
Overall Length		13,702 mm
Overall Canopy Width		8,295 mm
Overall Canopy Height		6,603 mm
Overall Tire Width		7,605 mm
Overall Height Body Raised		13,878 mm

Table 18-2: Mining Vehicle Repair Bay Specifications


Specification		793F Haul Truck
Bay Doors		10.0 m x 7.5 m
Bay Size		21.0 m x 18.0 m
Overhead Crane		50 t / 15 t
Crane Hook Height		14.0 m

The mine garage will have a total of six (6) maintenance bays, including two (2) bays for auxiliary vehicles and one (1) bay dedicated for welding. The bays will be aligned in a row with one (1) 50-tonne overhead crane servicing all bays. The building will include a 1,400 m² warehouse and a mechanical workshop that will also serve as a maintenance area for small vehicles. The workshop will be equipped with a 10-tonne overhead crane. The facility will also be equipped with a centralized lube distribution system for oils, grease and other fluids which will feed the lube stations at every bay. Every lube station will also have compressed air and service water outlets.

The truck wash facility will be located approximately 80 m south of the mine garage. The truck wash system will have mud-settling basins, a skimmer for oil and grease removal, and a water filtration system for continuous recycling of wash water. As a result, the filtration system will only require minimal make-up water (5-10%) to prevent mineral build-up. The building will also include a

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tire-change bay, designed to allow tire changes from inside the building on only one side of the vehicle. The other side will be accessible from outside the building through two (2) garage doors. Both the mine garage and truck wash buildings have been designed in a rectangular shape with an inclined roof to make them suitable for a pre-engineered structure. The structures are supported by shallow foundations founded on native soils.

18.4

General Offices and Assay Laboratory

All office facilities will be prefabricated type buildings made up of 12’ x 60’ modules. The buildings will be erected on concrete blocks with adjustable trestles founded on engineered fill. There will be three (3) main buildings: the main administration, the mine office and dry, and the plant office. The office requirements for each building are based on the staffing plans presented in Chapter 16. Table 18-3 shows a summary of the office staff requirements and office allocation.

Table 18-3: Staff Requirements


Total Requirements	Peak Personnel # Total		Personnel On Site # per Shift		Office Personnel Requirements
Open Pit Mining		299		112		33
Underground Mining		187		83		10
Process Plant		89		49		11
General and Administrative		26		20		26

Total:		601		264		80

18.4.1

Main Administration Building

The main administration building will be located at the entrance of the mine site and will house administration and safety/security staff only. The office design has ten (10) closed offices and space available for open workstations. There will also be a small cafeteria and three (3) meeting rooms.

18.4.2

Mine Office and Dry

The mine office will be located next to the truck shop and will house the mine, maintenance and engineering office staff. The building will also have dry facilities with lockers. There will be one (1)

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dry changing room for men and one (1) for women. The dry rooms will consist of lockers and sanitary installations designed to accommodate a varying men/women ratio of 90/10% to 75/25%, to allow for possible variations in staff. The locker requirements are doubled to include a dirty side locker and a clean side locker for each employee. The facility will also have meeting rooms, open areas for drawing reviews, a 136-seat lunch area which can also serve as a large conference/training room and a muster room on the ground floor for daily shift meetings.

18.4.3

Plant Office

The process plant office will be located on the west side of the process building between the leach tanks and the pre-leach thickener and will be connected to the main building via a short corridor. The building will house the process operations/maintenance office staff, and a dry facility. The building will also have a small lunch room and a 45-seat conference room.

18.5

Parking Area

Parking for employee vehicles and other service vehicles will be located in proximity to the process plant building and will have capacity for 150 vehicles. The parking area will also have space to accommodate two (2) passenger buses.

18.6

Assay Lab

The assay lab is located next to the parking area and is expected to handle approximately 200 samples per day and will have 17 employees. The assay lab includes a sample preparation area, wet laboratory area, fire assay, balance room, instrumentation room, an environmental lab, two (2) offices, a lunch room and two (2) washrooms.

18.7

Fuel Storage and Dispensing

The fuel island will be located close to the crusher on the main haul road to the plant and facilities. The tank farm will be located outside the blast radius of the pit, on the plant site road between the crusher and stock pile. The fuel island will have a diesel fuel pump station for mining vehicles and a containerized lube top-off system for oil, grease, windshield washing fluid and Extended Life

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Coolants (“ELC”). The fuel storage will be in seven (7) 80,000 L double-walled tanks. The total capacity is 560,000 L, which represents an autonomy of approximately six (6) days (95,000 L/day fleet consumption). Tanks will be prefabricated and delivered to site on skids with piping and pre-installed valve racks.

Diesel and gasoline will be made available for the light vehicle fleet which is expected to be mostly diesel pickup trucks. This light vehicle fueling station will be at a separate location from the fuel island that is required for the mining vehicles. The light vehicle fueling station will consist of horizontal double-walled tanks, equipped with all the required pumps and distribution equipment.

18.8

Explosive Plant and Storage

A platform for an explosive plant and storage depot will be built at a safe distance, north of the process plant area. These buildings will be supplied and constructed by the explosive supplier.

18.9

Electrical and Communication

18.9.1

Tie-Point Switching Station, Power Line, Main Substation, and Site Electrical Distribution

The total power demand of the Project was determined to be approximately 56.9 MW based on the estimated connected load, running load and running power.

Table 18-4 shows the power demand breakdown by sector. The power demand was calculated using an average efficiency factor, load factor and diversity factor. Specific numbers were used for the largest loads (primary crusher, SAG mill, ball mill and pebble crusher motors).

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Table 18-4: Estimated Total Project Power Demand


Area	Power Demand (MW)
Processing Plant¹		37.2
Site Infrastructure (including reclaim water barge)		5.9
Open Pit Mine		3.6
Underground Mine²		4.0
Network Loss (2%)		1.0
Power Demand Subtotal		51.7
Security Factor (10%)		5.2

Total Power Demand		56.9

¹	Excluding network loss.

²	Golder Associates provided a running load of 6.6 MW. Appropriate factors were applied to achieve the estimated power demand.

Electricity will be supplied to the site by a new 16.7 km long 230 kV power line, to be built and subsequently connected to the region’s existing 230 kV Hydro One line connecting Fort Frances and Kenora. For interconnection purposes, a 230 kV tie-point switching station is required.

A 115 kV supply from Hydro One could also potentially supply the plant as an alternative to the 230 kV connection. At the 230 kV interconnection point with Hydro One, there is also a 115 kV line present. The 115 kV supply option may be technically feasible and could be less expensive if capacity is available at the time the mine goes into service. However, this option would generate more line losses and would be less reliable than a 230 kV supply and would not be as flexible with regard to future expansion or configuration changes in the mine, such as additional milling capacity or in-pit crushing and conveying (“IPCC”). Considering the above points, the 230 kV line is considered to be the most practical option and is used as the basis for this Feasibility Study.

The main 230–27.6 kV substation will be located near the concentrator building where the large electric loads are installed. Both dual-pinion SAG and ball mills are equipped with one (1) active front-end Variable Frequency Drive (“VFD”) with an installed power of 10,000 HP per motor. The mills will be fed directly from the 27.6 kV switchgear. Two (2) main 230-27.6 kV, 60/80/100 MVA transformers will be used for a combined firm power of 100 MVA.

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This option provides ample room for a future expansion, while at the same time mitigating the risk of long downtimes due to transformer failure. A 36 kV Gas Insulated Switchgear (“GIS”), complete with electrical protection devices is included.

In addition to the mills, there will be several induction motors (“SCIM”), with powers ranging from 0.5 to 2,000 HP, which will require a 4.16 kV or a 600 V feed; therefore, for the 4.16 kV motors, three (3) 27.6-4.16 kV power transformers and two (2) 5 kV switchgears are required. For the 600 V motors, four (4) 27,600-600 V unit substations and one (1) 4,160-600 V unit substation are required.

The open pit mine will require electrical equipment that calls for 7.2 kV trailing cables for the hydraulic shovel. The mining electrical network will provide power at a voltage of 27.6 kV with one (1) 27.6-7.2 kV, 7.5 MVA substation. This will be used to step down the power near the mine equipment. The station will be optimally placed near the pit in order to minimize the 7.2 kV mine network and shorten the total length of the 7.2 kV trailing cables that are exposed to potential damage by mobile equipment and flying rocks during the blasts.

The electrical distribution to the site infrastructure will consist of a dedicated 27.6 kV OHL (overhead line) distribution network, equipped with 4/0 ACSR conductors.

The underground mine electrical distribution network will be developed and operated using the decline ramps. Electricity will be provided through the 27.6 kV mine OHL distribution network. A pole-mounted vacuum recloser (“VCR”) will be installed close to a 27.6 kV switchgear skid located at the primary crusher area. Mining power centers (27,000-600 V, 750 kVA or 1000 kVA) will be connected to this switchgear. After a few years, a second 27.6 kV switchgear skid will be required and installed at the ODM west vent shaft. This switchgear will also be connected to a pole-mounted VCR.

18.9.2

Emergency Power

Two (2) emergency power systems (4.16 kV and 600 V) are installed for the purpose of supplying the critical installations when the main power is lost. During a power outage, a Programmable

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Logic Control (“PLC”) will manage the critical loads. To achieve this task, the loads are regrouped under three (3) different categories; fixed loads (i.e., lighting, heating, leach tank agitator lube pumps), sequential loads (i.e., leach tank agitators, cyanide destruction tank agitators) and manually operated loads (i.e., sump pumps, rake mechanism, reactive heating).

At 4.16 kV, the critical loads are the tailings pumps, heat exchanger feed pumps, leach tank agitators and cyanide destruction tank agitators.

At 600 V, the main critical loads are as follows: process water pumps, service air compressor, tailing pumps, pre-leach thickener underflow pumps, CIP tank pump screens, pre-detox thickener underflow pumps, sump pumps and gland seal pumps.

A summary of the emergency power requirements for the process plant is presented in Table 18-5.

Table 18-5: Emergency Power Requirements


Location	MW		Voltage
Process Plant		1.7		4,160
Process Plant		1.3		600

The proposed emergency generator fleet for the process buildings, the reclaim water (barge) and the truck shop are presented in Table 18-6.

Table 18-6: Proposed Emergency Generators


Location	MW		Voltage
Crusher Building		0.25		600
Concentrator Building		1.50		600
Concentrator Building		2.50		4,160
Reclaim Water TMA		1.50		600
Truck Shop		1.50		600

The 4.16 kV emergency power supply of the process plant is a centralized system connected to the main 4.16 kV switchgear. However, generator starting is realized by a free-standing control panel. The system controls include a generator demand priority control function to automatically match the on-line generator capacity to the loads. The process plant 600 V emergency power

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supply is similar to the 4.16 kV systems, except that it is connected to dedicated switchgear that incorporates all the controls (starting and synchronizing).

18.9.3

Communication

A combined fibre optic self-healing loop backbone will interconnect all areas. This line will use the same poles as the 27.6 kV overhead distribution lines and can transmit voice, video and data on the following networks:

•

Telemetry, data acquisition, and control between the process plant and exterior process equipment;

•

Computer network between all departments;

•

Local telephone services; and

•

Computer network for maintenance on all electrical equipment data.

18.10

Tailings Management

The means of tailings deposition and storage is a key element for the operability and long-term closure strategy for the RRGP site. The TMA site was selected as it is in close proximity to the process plant and mine for tailings pumping. For dam construction considerations, it has relatively good topographic containment and is suitable for the enclosure of potentially acid generating PAG tailings.

The TMA covers an area of approximately 765 ha (excluding associated external ponds and infrastructure) and provides adequate storage capacity for the approximately 82 Mm³ of tailings anticipated to be produced over the projected mine life, based on an average deposited tailings dry density of 1.4 t/m³. The facility is bounded by natural topography (high ground) in the north-east and by impoundment dams along the remaining perimeter and has the potential for expansion if additional mineral resources are identified through ongoing exploration.

An overview of the TMA design is provided in the following sections. The design basis and analyses are provided in AMEC (2013c).

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18.10.1

Tailings Deposition Plan

Tailings will be deposited from spigots located along the inside perimeter of the TMA dams in order to develop a tailings beach in front of the dam. Perimeter discharge is a standard practice which enhances dam stability by keeping the pond away from the surrounding dams. The pond will be maintained within the central northern part of the basin to allow easy reclaim of water by barge and pump.

After approximately Year 9 (2024), the spigots will require extension inwards late in the mine life to fill the basin efficiently and minimize the volume of water required for the closure water cover over the tailings.

The starter dam has a crest elevation of 366.5 masl (~10.5 m in height) and can contain approximately two (2) years of tailings with 2 m of freeboard to the crest. The dams will be raised sequentially over the life of the mine to the maximum dam height of 23.5 m to contain the design tonnage.

Table 18-7 summarizes the design parameters for sizing the starter and ultimate dams.

Table 18-7: Tailings Dam Sizing


Parameter (Dike Structures)		Tailings Management Area
Starter Dam (2 years)
Dam crest elevation (masl)		366.5
Maximum height		10.5 m
Struck-level storage with 2 m freeboard to the crest		10,400,000 m³
Ultimate Dam
Dam crest elevation (masl)		379.5
Maximum height		23.5 m
Struck-level storage with 2 m freeboard to the crest		82,200,000 m³

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18.10.2

Dam Design

Dams, in general, have a very high hazard classification due to the potential environmental and economic consequences, in the event of a potential failure, and have therefore been designed for the most severe flood and earthquake criteria. The design criteria adopted are:

•

The site has low-to-moderate seismic risk with a 0.096 g horizontal peak ground acceleration for a 10,000-year return earthquake;

•

Required minimum FOS values for the design slopes are:

•

Short term, end of construction with induced pore pressures: 1.3

•

Long term, when excess pore pressures have fully dissipated: 1.5

•

Rapid draw down (of the Water Management Pond slope): 1.2

•

Worst case, with potentially slicken-sided upper varved clay: 1.0

•

Pseudo-static loading with a seismic coefficient of 50% of the peak ground acceleration (“PGA”): 1.0

The design dam slopes are 4H:1V for heights up to 15 m, which includes approximately 70% of the dam length. The higher south dam section requires a 6H:1V slope and an appropriately sized mine rock toe berm to reduce the overall slope to 6.5H:1V for stability. The slopes are designed to transition between the two (2) sections.

Stability analyses of the TMA dam slopes were carried out for critical sections using the limit equilibrium method for subsurface soil stratigraphy determined from the geotechnical investigations. The design and construction follow the observational design approach, considering the presence of thick, highly plastic clay and varved glacio-lacustrine units. Engineering judgement was used to select appropriate assumptions regarding excess pore pressures generated in the foundation soils and dam fill during construction that could govern stability. Dam instrumentation and monitoring will be required to be in place during the construction and operation phases. Details on the slope stability analyses are provided in AMEC (2013c).

The primary dam construction materials are select clay (overburden) and NPAG mine rock from open pit development which are available primarily during the pre-production and early years of

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operations. The majority of the TMA dam construction occurs during the first five (5) years of mine life.

The typical cross sections of the tailings dam are comprised of four (4) primary zones, as shown on Figure 18-1 and Figure 18-2. The legend is presented in Table 18-8:

•

Zone 1 is select overburden from open pit development that is placed in lifts with controlled compaction to act as the water retaining element of the dam;

•

Zone 2 is random overburden placed and compacted by the mine fleet to buttress the core and create the relatively shallow slopes dictated by the foundation soils;

•

Zone 3 is a NPAG mine rock toe for stability (South Dam only); and

•

Zone 4 is select or processed sand filter/drain that serves to prevent against cracks or construction defects from negatively impacting the performance of the dam.

Depending on the material balance Zone 2 of the dam could be constructed as a sandwich of overburden and mine rock, if suitable transitions between varying fill types are provided in critical areas.

18.10.3

Construction and Operational Considerations

The geometry of the dams allows for a significant portion of the construction material to be placed using haul trucks from the mine fleet. The dam core, filter, drain and erosion protection zones will be constructed by a qualified earthworks contractor.

TMA construction will follow the observational approach, which approaches the design on the basis of the expected conditions (i.e., excess pore pressures generated in the foundation), while accommodating a realistic worst case in the form of contingency plans that can be realistically implemented, if the observed conditions (i.e., measured pore pressures of displacements) indicate the need. Remedial works such as flattening the overall slope or increasing the size of the toe stabilization berm could be implemented within an appropriate timeframe.

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Overburden zones suitable for dam construction were delineated in the block model using trafficability criteria inferred from in situ water content (AMEC, 2012a). The mine plan and dam construction schedule are coordinated such that enough suitable overburden is available for dam construction in any given period. Further detail on the overburden sampling and characterization are provided in AMEC (2013a). A clay test embankment has been planned to guide the development of the overburden fill compaction specifications (AMEC, 2013d).

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Figure 18-1: Typical Cross Section - TMA South Dam Section (see Table 18-8)

Figure 18-2: Typical Cross Section – TMA West and North Sections

(Refer to Table 18-8 for the construction fill material descriptions)

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Table 18-8: Construction Fill Materials (Figures 18-4 and 18-5)


Construction Fill Materials
1		Core – select clay
2		Shell – random fill
3		Downstream Shell – clean mine rock
4		Filter – Sand
5		Transition – processed rock
6		Road Surface – sand & gravel
7		U/S Erosion Protection – cobbles & boulders
8		D/S Erosion Protection – cobbles
9		Bedding – sand & gravel
10		Armour Stone
11		Rip Rap

18.10.4

Site Water Management

The primary objectives of the water management system are to:

•

Generate a reliable water source for mill operations and ancillary uses;

•

Dewater the open pit and underground mine workings to ensure worker safety and operability;

•

Provide for general site drainage; and

•

Optimize the quantity and quality of site effluents released to the environment.

The system relies primarily on recycling water from various constructed ponds for mill process water in order to minimize the volume of fresh water to be taken from local watercourses. The system has been designed to ensure a reliable water supply at all times of the year and to allow for contingencies, such as dry years. The system includes five (5) constructed ponds for water management, in addition to one (1) temporary and two (2) permanent sediment control ponds, and one (1) direct freshwater source for operations (West Creek Pond). A constructed wetland with four (4) constructed ponds is to be built downstream of the TMA and may act as part of the site effluent treatment system. The lcoations of the water management infrastructure are shown in the general site drawings (Appendix H).

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Details of the water balance modelling and the overall site water management can be found in AMEC (2013e).

18.10.5

Water Management Structures

A brief description of each of the primary water management ponds required for the site water balance is provided below. The preliminary design characteristics of each pond are summarized in Table 18-8.

The MRP has been sized to operate based on the largest monthly pond volume for the 20-year wet annual precipitation conditions on an ultimate footprint of the east mine rock stockpile and open pit prior to the environmental design flood (“EDF”; a 24 hour/100-year return period storm event). Approximately 60% of the mill make-up water is provided from the mine rock pond. This rate was selected to ensure the pond can be kept in balance year over year in mean annual precipitation conditions. Regulation of water recycle to the process plant will ensure there is adequate storage available to contain the EDF with no discharge to the environment.

An area of approximately 375 ha reports to the stockpile pond. The stockpile pond emergency spillway is a rectangular spillway with an invert elevation of 365.5 masl with a dam crest elevation of 369 masl. The stockpile pond emergency spillway will have a width of 5 m with a design flow 141 m³/s and is only anticipated to be used during flood events greater than the EDF. The maximum storage capacity of the stockpile pond adjacent to the process plant is only about 80,000 m³, as it impinges on the toe of the ore stockpile. This pond receives runoff from the low grade ore and a portion of the NPAG mine rock stockpile. Water levels in the pond will be balanced by pumping excess water to the mine rock pond.

The WCP supplies all the fresh water to the process plant, as well as for site dust control from a watershed of 808 ha. An emergency spillway feeds into the West Creek diversion. The dam has been designed to hold back enough water to provide fresh water to the process plant for about four (4) winter months (about 216,000 m³) with an allowance for ice cover. The remainder of the water overflows to the environment through the West Creek diversion.

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A pond is expected to be developed at the head of the Clark Creek diversion (Clark Creek Pond) to facilitate the diversion into the lower reach of Clark Creek.

The TMA Pond (“TMAP”) is internal to the TMA and provides a relatively large volume of water for water supply to the process plant in extreme dry events. The TMAP has significant available storage capacity and can store excess water during wet events. For preliminary purposes, the TMA was sized to contain a minimum of 3.5 Mm³ of accessible water prior to winter. The TMA has ample capacity to contain the EDF event.

The Water Management Pond (“WMP”) receives the decant flows from the TMA for additional aging (greater than one (1) month proposed) and has a catchment area of 109 ha. It was sized for the wettest month of the 20-year annual wet conditions and will contain the EDF. The dam crest of the water management pond is 373.0 masl. The water supply required for start-up will be developed by constructing the water management pond in a single stage early in the project development. Further details on the WMP are provided in AMEC (2013e).

The Water Discharge Pond (“WDP”) receives decanted water from the water management pond and runoff from the local catchment area (100 ha) and decants low flows to the constructed wetland. The emergency spillway invert is at 354.3 masl and the dam crest elevation is 355.2 masl. Flows in excess of the wetland capacity will be directed to the Pinewood River to prevent damage to the constructed wetland due to high flows.

A constructed wetland is proposed to be established downstream of the water discharge pond within the Loslo Creek valley, upstream of the Pinewood River. The constructed wetland has been designed to take advantage of the natural topography present and support the additional passive treatment of a limited volume of discharge from the water discharge pond for nutrient and metal removal. The limited volume discharge will help mitigate flow reduction concerns associated with the Pinewood River and site surface water capture.

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A flood protection berm/access road is also proposed south of the open pit/NPAG stockpile area, to ensure that even under extreme flood conditions such as the EDF, the Pinewood River will not overflow into the open pit. Should the Pinewood River spill into the pit, it would cause excessive erosion of the overburden slope or potentially flood the pit. The preliminary flood protection design includes a 2.24 m high berm (including 0.3 m freeboard) having 3H:1V slopes and length of approximately 3,600 m, situated approximately 120 m from the Pinewood River. An access road will be situated on top of the berm.

18.10.6

Runoff and Seepage Collection

Runoff and seepage collection are required from mine site facilities as per Federal requirements. Runoff and seepage collected from the plant site area will be pumped to the MRP, either directly from the SP and mill area external sumps, or indirectly in the case of any runoff and seepage that bypasses these facilities and enters the open pit directly. The majority of the PAG mine rock stockpile area will drain by gravity to the MRP, however, portions of the eastern and southeastern boundary of the east mine rock stockpile outside of the Clark Creek basin could require separate constructed collection systems, depending on the final stockpile design. PAG mine rock will be placed within that portion of the PAG mine rock stockpile that will drain by gravity to the MRP, such that any acid rock drainage associated with this rock would be captured by the MRP both during operations and following mine closure. All runoff and seepage captured by the MRP will be contained within the overall water management system and will not be discharged directly to the environment.

Runoff and seepage collected along much of the south perimeter of the TMA in ditches will be routed through ditches to the WDP. Ditching bordering the northwest and west margins of the TMA will report to a collection pond, with this water being: released directly to the environment if it is of suitable quality; pumped back to the WMP if water quality is not suitable for direct discharge to the environment; or maintained in the water management system to enhance the existing water inventory.

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Runoff and seepage from the overburden/NPAG stockpile will be collected by east and west perimeter ditches that report to one or more terminal collection ponds located along the south boundary of the stockpile. Runoff collected from the overburden stockpile could contain relatively high levels of suspended solids, and as a result, will require sufficient retention time for the solids to settle out of solution naturally, and for the water to meet applicable water quality standards for direct release to the environment. If necessary, flocculants could be used to aid the settling of suspended solids.

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Table 18-8: Summary of Rainy River Gold Project Ponds


Pond	Discharge Pumped (P) / Decant or Spillway (D) to	Water Requirement	Maximum Operating Pond (20-year wet year) (Mm³)		Environmental Design Flood				Dam
Pond	Discharge Pumped (P) / Decant or Spillway (D) to	Water Requirement			EDF Runoff (Mm³)		Pond Volume including the EDF (Mm³)		Crest Elevation (masl)		Average Height (m)		Length (m)		Operating Period
MRP	To process plant (P)	Process water		2.93		0.31		3.24		362.0		5.4		1,650	January to December
SP	To MRP (P)	Transfer		0.08		0.00		0.08		369.0		4.0		155	January to December
WCP	To process plant (P)	Process and other fresh water needs		0.20		N/A		N/A		364.9		4.0		450	January to December
TMAP	To WMP (D)	Decanting for discharge		5.57		0.97		6.45		379.5		—		—	June to August
WMP	To the environment (Pinewood River below McCallum Creek; (P) To WDP (D)	Process water, with excess discharged to the environment		6.64		0.13		6.77		373.0		6.7		3,750	October, November, March, April, May January, June to September, December
WDP	To Constructed Wetland (D)	Excess discharged to the environment		0.08		0.03		0.112		355.2		1.2		360	January, June to September, December


Notes:		The maximum operating pond volume represents the largest monthly pond volume 20-year wet year.

		EDF - Environmental Design Flood, taken as the 1:100 year 24 hour storm event for ponds collecting mine affected water.

		All elevations are based on preliminary pond capacity information and required confirmation.

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18.10.7

Process Plant Water Supply – Preparations for Start-up

The primary water reservoir to support process plant start-up will be the WMP which is located immediately adjacent to, and southwest of the TMA. Construction of the WMP is planned to start once regulatory approvals are obtained. It is currently assumed that the WMP will be constructed and be able to begin receiving water inflow by the start of the spring freshet in 2015.

For the initial start-up, water will be taken from the Pinewood River and stored in the WMP for future use, in addition to natural inflows. A water intake structure will be constructed on the Pinewood River downstream of McCallum Creek. This location was chosen because the Pinewood River catchment and flow increases substantively downstream of the two (2) major tributaries, Tait Creek and McCallum Creek.

It is assumed that up to 20% of the spring flow (April to June; or March in the event of an early spring thaw), and up to 15% of the river flow during the period of July through November, will be withdrawn from the Pinewood River to develop the Project’s water inventory in the WMP. This approach is consistent with other Ontario mining projects. Water would be taken from the Pinewood River in 2015 and 2016. Thereafter, it is envisioned that there would be no direct water taking from the Pinewood River, except possibly for contingency purposes.

The available water from the Pinewood River under the percentage flow restrictions described above is shown in Table 18-9 for average and low runoff conditions. If flows approaching or above mean annual flow conditions are encountered, the percentage taken from the river would be reduced, as there would be excess water available under these conditions.

Table 18-9: Water Availability from the Pinewood River below the McCallum Creek Inflow


Condition	Month (‘000 m³)																Total
Condition	Apr		May		Jun		Jul		Aug		Sep		Oct		Nov		Total
Mean		2,233		1,716		1,260		571		277		312		424		334		7,127
5^thP		756		581		426		193		94		106		144		113		2,412
10^thP		928		713		523		237		115		130		176		139		2,961
25^thP		1,306		1,003		736		334		162		182		248		195		4,167


Note:		Tabled values represent a 20% taking of the spring flow (Apr-Jun) and a 15% taking for other months; no winter (Dec-Mar) water taking is proposed. Percentile (P) values are calculated as annualized and not monthly percentiles.

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18.10.8

Process Plant Water Supply – Operations

At full production capacity the process plant will require an approximate water input of 20,500 m³/d. Process plant outputs will include approximately 21,000 m³/d of water discharged to the TMA with the tailings slurry, and 400 m³/d of water lost to evaporation in the mill. Process water for process plant operations will result from recycling water from the MRP, as well as from the TMAP, the WMP and the WCP. Under typical, average annual operating conditions, an estimated 11,221 m³/d would derive from the MRP; 1,647 m³/d would derive from the WCP; 7,481 m³/d would derive from the TMAP; and 600 m³/d would enter the process plant with the ore. Ample water storage is available in the WMP and the TMAP to provide process plant water during the winter months, or during prolonged summer/fall drought conditions.

With regard to water availability in the TMAP, a portion of the water contained in the process plant slurry discharged to the TMA would be retained in the pore space within the deposited tailings. This expected water loss into permanent storage has been calculated as an average 7,092 m³/d. This value is based on a specific gravity of 2.82 for the ore and a settled tailings solids density of 1.41 t/m³. The difference between the volume of water in the tailings slurry discharged to the TMA, and the volume of water permanently stored within the tailings solids void space, would become available water for recycle back to the process plant for on-going processing. Excess TMA water not needed for processing would be discharged to the WMP for aging and then to the environment as treated effluent, generally directly to the Pinewood River, although some quantity will also flow through the constructed wetland, situated west of the overburden and NPAG mine rock piles.

Other sources of water for recycling to the process plant include precipitation and runoff collected within the TMA, and runoff and mine water that would be routed through the MRP. Modeling indicates that once steady state conditions are achieved, mine water will need to be removed at a net rate of approximately 6,600 to 9,800 m³/d, in order to maintain a reasonably dry and safe working environment. These values allow for the return of a small portion of mine water to support the mining operations, such as cooling water needed to support mine drilling activities. This excess mine water will be pumped to the MRP and will become part of the water inventory. The MRP will also collect natural runoff and seepage from the PAG Mine Rock Stockpile area, as well as water pumped from the SP located northwest of the stockpile near the processing plant. Upstream areas

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of the Clark Creek watershed will be diverted away from the PAG Mine Rock Stockpile area by means of the Clark Creek Diversion.

The MRP will provide a direct process water feed to the process plant, and will be designed with a storage capacity of approximately 2.93 Mm³. As the PAG Mine Rock Stockpile will store potentially acid generating and non-potentially acid generating rock, the MRP will contain water with increased suspended solids, possibly low levels of dissolved metals dependent upon geochemical reaction rates and residual ammonia from the use of blasting agents. If required, the water from the MRP will be drawn down by decreasing the amount of water going to the process plant from the TMAP and increasing the draw from the MRP.

In addition to the recycled water which will make up approximately 92% of water use in the process plant after start-up, a small amount of freshwater will be used for specialized process plant functions such as pump gland seal uses, and for reagent mixing. A dedicated pond, the WCP, will be established in line with West Creek for that purpose. Excess water from the WCP can be routed into the plant and sent to the TMA, if required, or, will be allowed to flow directly to the Pinewood River through the West Creek Diversion Channel. The WCP will only contain natural, non-contact water, and therefore does not require further management or treatment prior to release. The West Creek diversion channel will be kept separate from the Constructed Wetland downstream of the TMA, so as not to mix the natural creek water with TMA effluent.

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19.

MARKET STUDIES AND CONTRACTS

19.1

Market Studies

Neither BBA, nor Rainy River Resources, has conducted a market study in relation to the gold and silver doré which will be produced by the Rainy River Gold Project. Gold and silver are freely traded commodities on the world market for which there is a steady demand from numerous buyers.

19.2

Contracts

There are no refining agreements or sales contracts currently in place that are relevant to this Technical Report.

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20.

ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

20.1

General Approach

As demonstrated through its Health, Safety, Environment and Sustainability Policy, Rainy River is committed to excellence in the management of health, safety, the environment and sustainability in the conduct of its operations. The Company’s objectives are to:

•

Ensure the health and safety of employees, contractors and visitors in the workplace;

•

Responsibly manage the impact that its mineral exploration and development operations may cause to the environment; and

•

Demonstrate its commitment to fostering sustainable development in the communities in which it operates.

20.2

Consultation Activities

20.2.1

Community and Government Communications

Rainy River is an active member of the local community with offices in both Emo and Thunder Bay that offer residents easily accessible locations to learn about the Rainy River Gold Project. Rainy River has engaged the local communities as well as local First Nations and Métis community members in its Project planning activities. Through meetings, site tours, and regular communications, Rainy River strives to ensure engagement with all members of the local communities. Project open houses have been held in various communities to discuss the Project and to help people understand the Project as it moves forward.

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The involvement of stakeholders will continue throughout the various project stages over time. Key stakeholders who have demonstrated an interest in the Rainy River Gold Project include the following categories:

•

Municipal Governments (Township of Chapple, Township of Morley, Town of Emo, Town of Fort Frances and Town of Rainy River);

•

Various business, organizations and non-governmental organizations;

•

General public;

•

Federal Government (Aboriginal Affairs and Northern Development Canada, Canadian Environmental Assessment Agency, Environment Canada, Fisheries and Oceans Canada, Health Canada, International Joint Commission (Canada - United States), Major Projects Management Office, Natural Resources Canada and Transport Canada); and

•

Provincial (Ontario) Government (Ministry of Aboriginal Affairs, Ministry of Agriculture, Food and Rural Affairs, Ministry of Economic Development and Trade, Ministry of Energy, Ministry of Health and Long-Term Care, Ministry of Infrastructure, Ministry of Labour, Ministry of Municipal Affairs and Housing, Ministry of Natural Resources, Ministry of Northern Development and Mines, Ministry of the Environment, Ministry of Tourism, Culture and Sport, Ministry of Transportation and Hydro One Networks Inc.).

20.2.2

Aboriginal Communications

The Rainy River Gold Project site does not have any facilities or activities proposed on First Nation reserve lands or lands under land claim. The closest reserve to the Rainy River Gold Project is Rainy Lake Reserve 17b, located east of the proposed transmission line connection point and upstream of the Project.

While Rainy River recognizes that the land on which the Project is being planned is heavily impacted by historic and ongoing logging and some limited cattle ranching operations, the Company feels that it is important that Aboriginal oral history for the Project area be properly documented and respected. Rainy River is currently supporting preparation of a Traditional Knowledge/Traditional Land Use studies to assess use of the local are by Aboriginal peoples.

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Rainy River has not yet been apprised by the Métis Nation of Ontario whether the proposed development could have a negative effect on the Métis way of life. Meetings to understand any specific concerns are underway.

In anticipation of making a production decision, Rainy River requested advice from the Provincial Crown in 2010 as to which Aboriginal groups should be engaged regarding development of the Rainy River Gold Project and potential impacts on Aboriginal or Treaty rights. Following from advice provided by the Crown at the time, Rainy River engaged nine (9) First Nations that could potentially be affected by the Rainy River Gold Project, along with the Sunset County Métis:

•

Anishinaabeg of Naongashiing (Big Island) First Nation;

•

Couchiching First Nation;

•

Lac La Croix First Nation;

•

Mishkosiminiziibiing (Big Grassy River) First Nation;

•

Mitaanjigamiing (Stanjikoming) First Nation;

•

Naicatchewenin First Nation;

•

Nigigoonsiminikaaning (Nicickousemenecaning) First Nation;

•

Rainy River First Nation; and

•

Seine River First Nation.

In May 2012, the Provincial government identified changes and expanded considerably the list of Aboriginal groups to be consulted or notified about mine development. The majority of which are located in the Lake of the Woods area, a considerable distance from the Rainy River Gold Project site. Engagement with these additional First Nations has been initiated:

•

Naotkamegwanning (Whitefish Bay) First Nation;

•

Onigaming First Nation; and

•

Buffalo Point First Nation as directed by government agencies.

Rainy River is actively involving local Aboriginal groups in the Project planning and has negotiated comprehensive agreements with a number of the local First Nations to set protocols and commitments for ongoing involvement for the life of the project and community benefits that would, in part, help mitigate any potential adverse effects to Aboriginal or Treaty rights.

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Rainy River and the Fort Frances Chiefs Secretariat signed a Memorandum of Understanding in May 2010 which commits Rainy River Resources to informing the Fort Frances Chiefs Secretariat in advance about exploration proposals and schedules, and conducting exploration activities in an environmentally responsible manner. Employment and contracting opportunities were also part of the terms including the joint initiative to fund the First Nations Engagement Specialist position (for the Fort Frances Chiefs Secretariat). The Memorandum of Understanding also committed Rainy River Resources to developing and implementing a Participation Agreement (also known as an Impact and Benefits Agreement) that would include provisions for: mineral production support, consultation protocols, respect for Traditional Territories, training and employment, among other aspects. As a result of various community and leadership meetings conducted since the signing of the Memorandum of Understanding, a Participation Agreement was successfully developed with six of the First Nations and subsequently signed in March of 2012. A Participation Agreement Advisory Committee was formed comprising First Nations and Rainy River Resources representatives. The Participation Agreement Advisory Committee currently meets on a monthly basis to share information and ensure the successful implementation of the Agreement.

Rainy River also signed a Memorandum of Understanding on March 6, 2012 with the Big Grassy River First Nation, and is also engaging the Big Island First Nation, Onigaming First Nation and Naotkamegwanning (Whitefish Bay) First Nation. Rainy River has also been engaging with the Region One Métis Consultation Committee.

Key issues and interests raised by Aboriginal groups to-date are those related to: environmental management; employment and benefits; fisheries and wildlife; Project components and mining; traditional land use and culture; and water resources.

While the Rainy River Gold Project is located primarily on patented lands or lands that have been used for forestry and agricultural purposes for several generations, Rainy River is nonetheless endeavouring to engage its Aboriginal partners in the collection and documentation of Traditional Knowledge and Traditional Land Use. The Company feels that it is important to ensure that any traditional land uses be properly documented and used respectfully to avoid, or limit potential effects.

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20.3

Environmental Studies

20.3.1

Overview

Four (4) years of environmental baseline investigations have been completed in support of the Rainy River Gold Project. The description of the existing environment provided herein is a focused summary based on extensive baseline studies conducted to date for the Rainy River Gold Project. The intent of this section is to familiarize the reader with the local setting. Further detail, including copies of the baseline reports referenced herein, will be provided in the Environmental Assessment.

The objectives of the baseline studies are to characterize the natural (or biophysical) and human environment aspects of potentially impacted areas, as well as reference locations (such as upstream locations) where appropriate for comparison. Environmental baseline data (description of the existing environment):

•

Helps inform Project designs (for example, knowledge of rock characteristics assists in determining how best to handle and store the material);

•

Will allow an assessment to be made of likely Project environmental effects, including comparisons with established environmental guidelines, thresholds and limits, where applicable; and

•

Provide a reference for future environmental monitoring (that is, it allows a comparison to be made of pre-development and post development conditions).

Studies to-date have been completed using standard field protocols and scientific methodologies, to accurately document a real and temporal variability, and have considered the information needs of regulatory agencies for approval of previous Ontario mining projects. The baseline studies have included collection of site-specific information regarding the following general aspects (and others) as well as documentation of applicable published material:

•

Meteorological, air quality and noise data;

•

Vegetation mapping;

•

Field surveys of breeding birds, amphibians and mammals, including Species at Risk;

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•

Aquatic habitat mapping, fish community assessment and fish flesh analysis;

•

Sediment quality and determination of benthic invertebrate assemblages;

•

Surface water flows and quality;

•

Hydrogeological investigations and groundwater quality;

•

Geochemistry of existing mineral wastes and future deposits;

•

Socio-economic characterization and land use; and

•

Heritage resources and archaeological surveys.

Traditional Knowledge and Traditional Land Use data are also being collected as part of the ongoing Aboriginal consultation program.

Environmental studies summarized herein are based on the documents listed below. Certain aspects are still ongoing:

•

KCB (2011): Rainy River Gold Project - Baseline Report, 2008 - 2010;

•

AMEC (2012): Rainy River Gold Project - Climate, Air Quality and Sound Baseline Study (report in progress);

•

AMEC (2012): Rainy River Gold Project - 2011 Aquatic Resources Baseline Study;

•

AMEC (2012): Rainy River Gold Project - 2011 Site Noise Monitoring Report;

•

AMEC (2012): Rainy River Gold Project - Socio-Economic Baseline Study (report in progress);

•

AMEC (2012): Rainy River Gold Project - 2011 Terrestrial Resources Baseline Study;

•

AMEC (2012): Rainy River Gold Project - 2011 Species at Risk Report;

•

AMEC (2012): Interim Report, Metal Leaching and Acid Rock Drainage Characterization of Mine Rock, Geochemistry Report - Rainy River Gold Project;

•

AMEC (2012): Rainy River Gold Project - 2012 Terrestrial Resources Baseline Study (report in progress);

•

AMEC (2012): Rainy River Gold Project - 2012 Species at Risk Report (report in progress);

•

AMEC (2012): Rainy River Gold Project - 2012 Aquatic Resources Baseline Study (report in progress);

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•

AMEC (2012): Rainy River Gold Project - 2012 Site Sound Monitoring Report (report in progress);

•

Woodland Heritage Services (2012): Stage 1 Archaeological and Cultural Heritage Resource Assessment of the Rainy River Resources Advanced Exploration Project, northwest of Fort Frances, Rainy River District, Ontario (in progress); and

•

Woodland Heritage Services (2012): Stage 2 Archaeological and Cultural Heritage Resource Assessment of the Rainy River Gold Project, northwest of Fort Frances, Rainy River District, Ontario (in progress).

20.3.2

Climate, Air Quality and Sound

20.3.2.1

Climate

The nearest Environment Canada climate station to the Rainy River Gold Project site, for which long-term, current records are available, is located at Barwick, Ontario. This station is located 23 km southwest from the site, at coordinates 428807E and 5387043N, and has climate records dating back to 1978. Mean monthly temperatures range from a low of -15.9°C in January to a high of 18.8°C in July. The mean annual precipitation for Barwick is 694.7 mm, with 79.5% of this average value occurring as rain. June is typically the wettest month.

A dedicated climate station has been operating at the Project site since June of 2009. Temperature and precipitation results show strong agreement between the Project site and Barwick station precipitation records. Wind speeds for the Rainy River Gold Project climate station show average daily speeds ranging from 2 to 15 km/h, with maximum daily wind speeds ranging from 10 to 80 km/h. There was no overall dominant wind direction noted, but the strongest sustained prevailing winds tend to come from the northwest and the southeast.

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20.3.2.2

Air Quality

Background air quality at the Rainy River Gold Project area is expected to be good, given the absence of nearby large urban centres and industrial sources. Air quality in the Rainy River Gold Project area will, however, be influenced by long-range transport of air emissions from the south and also by natural sources, such as volatile organic emissions from vegetation and natural fires. The greatest potential local influence to air quality is increased particulate matter from traffic, logging/cattle ranching operations and drilling. For this reason, air quality baseline conditions were developed primarily based on published sources for most parameters.

Background air quality data were measured at two (2) stations proximal to the Rainy River Gold Project site to measure particulate matter, with stations located respectively 115 m and 20 m from Highway 600. Highway 600 (a gravel road) was considered to be principal source of local particulate material (road dust). All measured average particulate material concentrations were below applicable Federal and Provincial standards/criteria.

20.3.2.3

Sound

Ambient noise surveys have been conducted periodically at the Project site since 2009. Measured sound levels have been relatively consistent with background Energy Equivalent Continuous Sound Levels (“Leq levels”) generally ranging from about 40 to 50 A-weighted decibels for most sites, but with the more distant sites (sites >200 m from roads) showing more typical values in the 35 to 40 A-weighted decibel range. Differences between day and night time sound levels were not appreciable.

20.3.3

Physiography, Soils and Geology

20.3.3.1

Physiography

Terrain in the Project site area transitions from upland bedrock controlled lake areas to the northeast, to lower-lying to gently undulating terrain to the southwest. The Pinewood River system which drains most of the Project site area occupies a broad lacustrine plain. Lands in the

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immediate Project site vicinity are typically gently rolling to flat, with wetlands occurring in low-lying areas, and rounded bedrock outcrops and subcrops occurring in upland areas. Elevations increase to as much as 430 m above sea level in highlands northeast of the Project site, and decline to approximately 340 m above sea level in lower reaches of the Pinewood River valley southwest of the Rainy River Gold Project site. Maximum slopes in localized areas are typically in the order of 5 to 10%. Low-lying areas were inundated by glacial Lake Agassiz which left a variably thick veneer of lacustrine clay over much of the landscape.

The overburden sequence at Project site consists of discontinuous Labradorean Till, overlain by Keewatin Till, with the Keewatin Till typically being overlain and underlain by Lake Agassiz lacustrine deposits. Extensive surface peat deposits occur in many low-lying areas. Alluvial deposits occur in the creek and river valleys. The Labradorean Till consists mainly of silty sands and gravels, and forms localized aquifers, frequently kept under pressure by the overlying lower permeability Keewatin Till. The Keewatin Till is clay-rich and clast poor and is prevalent throughout the area except where it is disrupted by bedrock and subcrop zones. Average Keewatin Till thickness at the Rainy River Gold Project site is in the order of 20 to 25 m, whereas the underlying Labradorean Till is typically less than 5 m in thickness, and is discontinuous. Lake Agassiz lacustrine sediments, comprising clays, with minor silts and sands, typically occur above and below the Keewatin Till, but can also occur locally in more complex sequences. Overall overburden thicknesses can range up to 100 m in some places, but are typically in the order of 20 to 30 m in areas closer to the Project site, which are not disrupted by bedrock exposures. Peat deposits are typically <1.5 m in thickness but can be thicker and are widespread in low-lying areas.

20.3.3.2

Soils

Soils in the Project area are generally comprised of gray luvisols, gleysols, humisols, and rockland soils, with lesser expressions of podzolic and brunisolic soils. Gray luvisols are typically clay or clay/silt rich and imperfectly drained. Gleysols are poorly drained/frequently saturated, and in the Project site area generally consist of silt loams to more coarse textured soils. Humisols (organic soils) are associated with wetland systems. From a textural perspective the majority of Project site area soils consist of clay and clay loam soils, with lesser quantities of sandy clay loam, sandy

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loam, loam, silt loam and silty clay. Site specific investigations included 95 soil samples collected from 50 test pits.

The Project soils are overwhelmingly calcareous, except for the organic soils, due to the nature of the parent material. Organic soils are acidic. Cation exchange capacity tends to be relatively high because of the elevated organic and clay content of most soils present. Soil metal contents were typical of expected background soil conditions for the soil types present.

20.3.3.3

Geochemistry

Section 7 provides a description of local geology from a resource-based perspective.

Gold mineralization is associated with earlier sulphide formations consisting of pyrite, sphalerite, chalcopyrite and galena stringers and veins, and disseminated pyrite, together with later formed quartz-pyrite-chalcopyrite veins and veinlets.

Static geochemical testing for acid-base accounting, total metals, and shake flask analysis was conducted on 658 samples, 366 of which represent expected pit wastes consisting of 10 identified lithologies. The remaining 292 samples represent ore, or samples peripheral to the currently proposed open pit. An additional 20 samples were selected for kinetic testing, and 7 field trials are underway. Acid-base accounting testing was conducted using standard procedures including the Modified Sobek method for neutralization potential determination. An estimated 42% of the mine rock is considered to be potentially acid generating. Neutral metal leaching is not expected to be a concern for the Project. Kinetic testing indicates that onset times for acid rock drainage development are likely to be on the order of several years to decades depending upon the material type thereby providing sufficient time for implementation of appropriate reclamation methods.

Testing of simulated tailings material has also been conducted with samples from the metallurgical program. The tailings are potentially acid-generating, and will require management as such. Kinetic testing indicates that onset times for acid rock drainage of tailings materials are likely to be in the order of 27 to 41 years.

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20.3.4

Hydrology and Hydrogeology

20.3.4.1

Hydrology

Local drainage systems are characterized by numerous small creeks draining to the Pinewood River. The creeks generally originate in rocky uplands, but also frequently originate from or pass through headwater wetland systems. Hydrological systems to the northeast (upstream) of the Project site show an abrupt transition to larger lake systems in bedrock dominated terrain. This lake terrain is remote from proposed project development areas with the exception of the transmission line corridor which passes through this northeast area, but will not impact directly on any lakes.

Much of the Project area has been cleared over the years for both timber harvesting as well as through cattle ranching. Most of the natural drainage systems have been altered near the Rainy River Gold Project site through the development of agricultural drains and ongoing beaver activities. In addition to the Water Survey of Canada data, water level/flow data have been collected periodically from local creek systems.

As with all of northern Ontario, peak stream flows occur in the spring, with a secondary smaller peak flow in the fall. Low flows occur in the winter under ice cover, and also more variably depending on the year in the late summer or early autumn. The average annual runoff for the Project site area is approximately 200 mm, reflecting progressively drier conditions towards the western portion of the province. Groundwater base flow to area creeks is limited due to the prevailing clay substrates, such that, zero or near zero flows are experienced in both local creeks and the Pinewood River during late winter and late summer periods in some years.

20.3.4.2

Hydrogeology

The groundwater regime is governed by the overall structure and hydraulic properties of the overburden and bedrock sequences, and by the local topography and associated surface watercourses. A network of over 100 installations has been used to assess site area groundwater conditions, consisting of monitoring wells, test wells, drill holes, auger holes, mini-piezometers, and cone penetration test holes. AMEC has developed a hydrogeological model using the Modular

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Finite-Difference Groundwater Flow Model to estimate likely seepage rates of groundwater into a fully developed, open pit.

Based on model applications and sensitivity analyses, the predicted groundwater seepage rates into the open pit are expected to be in the order of 1,800 m³/d to 3,200 m³/d at full open pit development. The predicted drawdown cone from dewatering of the open pit is predicted to extend approximately 2 km to the west, 1 km to the east and 2.5 km to the south and north from the pit by the end of mining. No private wells will be located within the current estimated drawdown cone. Parts of the Pinewood River and several of its tributaries also lie within the projected drawdown cone. The impacts of a reduction in groundwater discharge to the Pinewood River are expected to be minor and difficult to measure, given that flow in these sections of the river and its tributaries are dominated by surface water runoff contributions and these features can often be dry. A monitoring network of monitoring wells and surface water level stations is proposed to confirm predicted impacts.

Additional groundwater modelling is underway to assess the potential effects of open pit dewatering on other area creeks and wetlands. Based on the high clay content of the Keewatin Till and the associated glacial Lake Agassiz sediments, these creek and wetland systems are expected to be perched, and not overly sensitive to open pit dewatering effects.

20.3.5

Surface Water, Sediment and Groundwater Quality

20.3.5.1

Surface Water

A total of 20 surface water sampling stations were established for the Rainy River Gold Project during the period of 2007 through 2010, with 14 of these stations still being active. Attempts have been made to position stations upstream and downstream of potential future developments within the limitations of the local drainage systems. Since June 2010, water quality samples have been collected at approximately monthly intervals to provide a full seasonal spectrum of data, with samples being analysed for a broad range of general parameters and metals. The analysis of general parameters includes: pH, conductivity, total suspended solids, ammonia, fluoride, nitrite, nitrate, total cyanide and E. coli. The inductively coupled plasma metal scan includes dissolved

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aluminum, and total metal concentrations for antimony, arsenic, boron, cadmium, chromium, cobalt, copper, iron, lead, mercury, molybdenum, nickel, thallium, uranium, vanadium and zinc.

Surface water quality in the area is generally quite good, with all parameters typically meeting applicable objectives/guidelines for the protection of aquatic life except for: iron, aluminum and phosphorus, which were commonly above their respective objectives; cadmium, copper and cobalt, which occasionally to commonly exceeded either Federal or Provincial objectives; and occasional to rare exceedances for arsenic, lead, nickel and zinc. Increased coliform levels were also noted at some stations, which may be related to area cattle operations and cattle foraging activities. It is not unusual for baseline water quality to exceed objectives/guidelines for various metals as a result of high suspended solids loadings in some samples, naturally elevated metal content in the local soil and rock, and because of seasonal ion concentration processes involving ice formation in winter and evaporative process in summer. Erodible clay/silt soils, cattle activity, and low creek base flow conditions (making them prone to ion concentration effects), are all contributing factors to observed water quality conditions.

20.3.5.2

Sediment

Sediment quality samples were collected from 2008 to 2012 from various upstream and downstream stations. Sample analyses included pH, particle size, chloride, bromide, sulphate, nitrite, nitrate, phosphate, total organic carbon, and a metals scan for aluminum, antimony, arsenic, barium, beryllium, cadmium, calcium, chromium, cobalt, copper, iron, lead, manganese, mercury, molybdenum, nickel, potassium, sulphur, selenium, silver, sodium, strontium, thallium, vanadium and zinc. Sediment quality is generally good with parameters generally found below Federal guideline values, and below Provincial sediment quality guideline lowest effect levels.

20.3.5.3

Groundwater

Groundwater quality samples were collected from monitoring well and drill holes near to the Project site periodically from 2007 to present. Samples were analyzed for pH, conductivity, total organic carbon, colour, chloride, fluoride, ammonia, nitrite, nitrate, biochemical oxygen demand, total and fecal coliforms, total cyanide, oil and grease, and metals, including: aluminum, antimony, arsenic, barium, beryllium, bismuth, boron, cadmium, calcium, chromium, cobalt, copper, iron, lead, lithium,

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magnesium, manganese, mercury, molybdenum, nickel, potassium, selenium, silver, sodium, strontium, tellurium, thallium, thorium, tin, titanium, tungsten, uranium, vanadium, zinc and zirconium. Data were compared with Federal and Provincial, objectives and guidelines for the protection of aquatic life, and drinking water standards, recognizing that these criteria are not directly applicable.

Groundwater baseline water quality is expected to reflect the naturally elevated metal content in the local soil and rock. Analytical results for Project site drill holes and wells showed that a number of groundwater samples exceeded objectives and guideline values for the protection of aquatic life for cobalt and iron, and that occasional exceedances of protection of aquatic life guidelines were noted for cadmium, copper, molybdenum, uranium, and zinc. Federal or Provincial drinking water guidelines were commonly exceeded for total dissolved solids, turbidity, iron and manganese; with occasional exceedances for barium, chloride and antimony.

20.3.6

Biological Environment - Existing Conditions

20.3.6.1

Aquatic Resources

Studies of fisheries and aquatic resources were carried out during 2008 through 2012. These studies included habitat assessments and fishing efforts focused on the Pinewood River system and its tributaries using a variety of fishing gear (electroshocking, seines, minnow traps, trap nets and gill nets). Fishing efforts were conducted in spring, summer and autumn, together with some late winter sampling efforts.

In the general vicinity of the Project area the Pinewood River shows typical widths of 10 to 15 m, with wider sections associated with beaver impoundments. Summer water depths are typically 0.9 to 1.7 m, with maximum summer water depths in the order of 2 m. Substrates consist of clays and silts, with some detritus. Gravel, rock or cobble substrates are almost nonexistent. Turbidity is high because of erosion of the clay and silt substrates, and agricultural drainage. Cattle wade into the river at a number of locations. Dissolved organic carbon levels are also elevated due to wetland drainage, further contributing to poor water column visibility. Beaver dams are frequent.

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The smaller creeks draining into the Pinewood River (Loslo, Marr, West, Clark, Tait and Blackhawk Creeks) typically exhibit summer widths of 0.5 to 3 m, except where they are impounded by beaver dams, with upper creek reaches being smaller, generally from <0.5 to 1.5 m and frequently exhibiting intermittent flow. Headwater areas of many of these tributary creek systems are associated with wetland systems. Beaver impoundments are frequent.

Large-bodied species (Northern Pike, Brown Bullhead and White Sucker) were found only in the Pinewood River, and not in the smaller tributaries, with the exception of White Sucker which were also found in Clark Creek. The distribution of Northern Pike in the Pinewood River appears to be limited to, or principally limited to areas downstream of the Blackhawk Road crossing area, possibly due to habitat limitations or more likely because of restrictions imposed by numerous beaver dams on the system. Northern Pike are reported to have been present in Pinewood Lake in the past, but have not been observed in several years. Local residents indicated that Walleye and Yellow Perch occur further downstream in the Pinewood River, but not in the general area of the Project site. Lake Sturgeon occur in the Rainy River, but according to local knowledge are not known to occur in the upper Pinewood River. Neither Walleye nor Lake Sturgeon have been observed in any of the aquatic surveys to date related to the Rainy River Gold Project, including downstream sections of the Pinewood River.

Small-bodied fish occur in greatest abundance in the smaller tributaries, possibly because of an absence of predaceous Pike in these tributaries. There are no salmonid (trout) species present. West Creek and Clark Creek are apparently used by local bait fishermen.

Benthos communities were assessed in the Pinewood River and in site area tributaries. Benthos metrics included densities, richness, evenness and the Bray Curtis Index of dissimilarity, in accordance with Environmental Effects Monitoring protocols. The benthic community generally exhibits a low-to-moderate abundance, and is dominated by crustaceans and chironomids, most likely because of prevalence of clay-silt substrates, which provide limited habitat conditions.

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20.3.6.2

Vegetation Communities

The Rainy River Gold Project and environs occur within the Agassiz Clay Plain Ecoregion, which extends from Lake of the Woods in the west to Fort Frances in the east, and from the United States border northward. The Pinewood River watershed is dominated by mixed Poplar and Black Spruce forests, and by non-forested areas (mainly agricultural lands), together with wetlands. The local area shows an even greater preponderance of mixed poplar forests, which occupy greater than 50% of the landscape, together with wetlands and agricultural lands. Wetlands are comprised mainly of treed and open fens, together with wetland thickets and marsh areas. Agricultural lands are mainly pasture and hay fields. Poplar forests comprised principally of Trembling Aspen are indicative of disturbed lands as Trembling Aspen are a successional species in Ontario.

Only two (2) Provincially rare species have been identified in the local area: New England Violet and Field Sedge. Muskroot, another rare species, has been identified as being present historically. There are no specific approvals related to these species.

20.3.6.3

Wildlife

In-field wildlife surveys were carried out for the Rainy River Gold Project and its immediate environs. These surveys were focused principally on birds and to a lesser extent on mammals, amphibians and dragonflies/damselflies. A proactive focus was placed on Species at Risk assessment and permitting planning by Rainy River Resources with the Ministry of Natural Resources.

Focused surveys have been conducted on forest birds, breeding birds, owls, marsh birds and waterfowl species, Sharp-tailed Grouse, nocturnal avian species (Whip-poor-will, Common Nighthawk and owls), raptors and amphibians, using established survey protocols. The relatively high avian species diversity present in the area reflects the mosaic of principal habitats in the areas (forest, wetlands, fields and shrublands), and the transitional (or near transitional) position of the study area relative to the Great Lakes, Boreal and Prairie regions.

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Twenty-two mammal species have been identified in the Rainy River Gold Project environs through direct observation, trapping records or sign. A number of other small mammal species are also likely to be present as no specific efforts were made to trap small mammals. Three (3) commercial trap lines overlap with the local area. Fur returns for these trap lines for the period of 1993 through 2008 indicated that Beaver, American Marten, Red Fox, Otter, Fisher and Mink are the most frequently and valued furbearers taken.

Amphibian and reptile surveys identified eight (8) frog species and three (3) reptile species (Eastern Garter Snake, Western Painted Turtle and Snapping Turtle). No salamander species were observed. Twelve species of dragonflies/damselflies were observed, or are known to occur, in or adjacent to the Rainy River Gold Project, three (3) of which are Provincially rare (Horned Clubtail, Arrowhead Spiketail and Green-faced Clubtail) and do not require special permit or authorization considerations.

20.3.6.4

Species at Risk and Critical Habitat

The Species at Risk known to occur in the Rainy River Gold Project environs are listed in Table 20-1. Rainy River Resources has been working closely with the Ministry of Natural Resources since June of 2010 in support of meeting permitting requirements of the Ontario Endangered Species Act. Permits are only expected to be required for a limited number of species.

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Table 20-1: Species at Risk Known to be Present in the Rainy River Gold Project Environs


Species	Conservation Status
Common Name	SARO	SARA	COSEWIC
Birds
Barn Swallow	T	—	—
Bobolink	T	—	T
Whip-poor-will	T	T	T
American White Pelican	T	NAR	NAR
Bald Eagle	SC	NAR	SC
Canada Warbler	SC	T	T
Common Nighthawk	SC	T	T
Golden-winged Warbler	SC	T	T
Olive-sided Flycatcher	SC	T	T
Peregrine Falcon (migrant)	T	SC	T
Red-headed Woodpecker	SC	T	T
Short-eared Owl	SC	SC	SC
Mammals
Little Brown Myotis (bat)	T	T	T
Northern Myotis (bat)	T	T	T
Reptiles
Snapping Turtle	SC	SC	SC


Notes:		E – Endangered, NAR – Not at Risk, SC – Special Concern, T – Threatened.

SARA		Species at Risk Act and is the Federal Status. Rankings are provided by the Committee on the Status of Endangered Wildlife in Canada (COSEWIC).

SARO		Species at Risk in Ontario and is the Provincial Status. Rankings are provided by the Committee on the Status of Species at Risk in Ontario (COSSARO).

20.3.7

Human Environment

20.3.7.1

Population and Demographics

The population of the Rainy River District was 17,912 in the 2011 Census, a decline of 1.8% from the 2006 Census, itself a 2.5% decline from the 2001 Census. Approximately 55% of the Rainy River District’s population resides in the largest urban centre, Fort Frances that has a population

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of 7,952. Excluding the smaller hamlets, Emo is the closest community to the Rainy River Gold Project site with a population of 1,252. The trend of population decline is expected to continue in the region over the long term (Ministry of Finance 2009), due, in part, to loss of employment in the forestry sector.

The region has approximately equal representation of males and females with the youngest and oldest age cohorts being higher than the Provincial averages. Overall, the median age is higher than the province with a median age of 43.2 and even older in the rural areas (unorganized) of the District.

20.3.7.2

Regional Economy

The regional economy in the Rainy River District is primarily supported by the forestry sector with three of the ten major employers involved in forestry manufacturing. The remainder of the major employers are in public and retail services (health, education, municipal government). In 2006, the participation in the labour force for the Rainy River district was 64.2%. This rate is slightly lower than for the province (67.1%) with a significantly higher unemployment rate (7.9%) compared with the province (6.4%). The Rainy River district had a larger workforce share employed in occupations unique to primary industry and trades, transport and equipment operates and related occupations in 2006, compared with Ontario as a whole. This suggests that the workforce is well positioned to service the Rainy River Gold Project.

20.3.7.3

Community Infrastructure and Services

Given that the region has experienced population declines, service capacity may be able to handle additional demands which could be experienced by these communities in the event of population increases either temporarily during the construction phase or permanently in operations phase of the Rainy River Gold Project. The available information suggests that there could be some near-term capacity challenges for housing and accommodation, as well as for cellular communications services which Rainy River Resources is currently investigating.

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The region is very well serviced and accessible from Highways 71, 11 and 600. The Canadian National Railway runs east-west through the region and within 40 km of the Project site with links to Winnipeg (Manitoba), Thunder Bay (Ontario) and Duluth (Minnesota). Fort Frances has regular commercial air service.

20.4

Environmental Sensitivities

There are a number of Species at Risk known or expected to occur within the Rainy River Gold Project footprint or environs. These include three (3) bird species and two (2) bat species listed as “Threatened” provincially (Whip-poor-will, Bobolink and Barn Swallow; Little Brown Myotis and Northern Myotis). Based on information currently available, Provincial Species at Risk Permits, pursuant to requirements of the Endangered Species Act, are likely to be required for Whip-poor-will and Bobolink. It is anticipated that the required Provincial Species at Risk Permits for these two (2) species, and possibly others if applicable, will be issued for the Rainy River Gold Project to proceed, but avoidance of critical Species at Risk habitat is one the mitigation measures required which has already been incorporated into the Feasibility Study design to a level practical.

There are no Federal lands within the Project footprint and therefore, Federal Species at Risk Permits will therefore not be required.

None of the forested or wetland vegetation communities investigated are provincially rare and there are no endangered species within the regional Project area.

There are no Areas of Natural Scientific and Interest, Environmentally Sensitive Areas or Provincially Significant Wetlands within the Project area.

20.5

Regulatory Context

20.5.1

Current Regulatory Status

Exploration has been conducted at the Project site since 2007. Based on AMEC’s site inspections to-date, the exploration has been completed in an appropriate manner from an environmental

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perspective. AMEC understands that the site is fully compliant with existing environmental approvals and has conducted the exploration with due regard to environmental considerations.

20.5.2

Environmental Approvals Required for Proposed Operations

20.5.2.1

Environmental Assessments

Most mining projects in Canada are reviewed under one or more Environmental Assessment (the “EA”) processes. The Environmental Assessment process is a means of project review whereby project design choices, environmental impacts and proposed mitigation measures are compared and reviewed to determine how best to proceed through the environmental approvals and permitting stages. Entities involved in the review process normally include government agencies, Municipalities, Aboriginal groups, various interested parties and the general public.

The Rainy River Gold Project requires completion of a Federal Environmental Assessment, pursuant to the Canadian Environmental Assessment Act, 2012. The Federal “Regulation Designating Physical Activities” identifies the physical activities that constitute the designated projects that could require an EA. Section 15(d) of the Regulation identifies one of the designated projects as: “the construction, operation, decommissioning and abandonment of a gold mine, other than a placer mine, with an ore production capacity of 600 tonnes per day or more”.

Rainy River submitted a Project Description to the Canadian Environmental Assessment Agency in August 2012, which was accepted. Based on the Project Description, the Canadian Environmental Assessment Agency confirmed that a Federal Environmental Assessment is required. The Environmental Impact Statement Guidelines, which identify the scope of the EA required for the Project, were issued on December 18, 2012.

In addition to needing to meet Federal Environmental Assessment requirements, Rainy River entered into a Voluntary Agreement with the Ontario Ministry of the Environment on May 4, 2012, to conduct a Provincial EA for the Rainy River Gold Project that will meet the requirements of the Ontario Environmental Assessment Act. Rainy River entered into the Voluntary Agreement in order to facilitate meeting the Provincial EA requirements to allow issuance of Provincial approvals to construct the mine. Several aspects of the Project were anticipated to require completion of

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Provincial Environmental Assessment process(es). A single Provincial process coordinated with the required Federal EA process, was selected as the best approach to meet those (or other) needs, as per the following:

•

A 230 kV transmission line of approximately 20 km length;

•

Diesel generation of between 1 and 5 MW generation;

•

Disposition of Crown resources, potentially related Crown lands (such as work on streambeds/shorelands) and effects on Species at Risk; and

•

Realignment of gravel-surfaced Highway 600 to avoid potential land use conflicts.

In parallel with the submission of the Project Description to begin the Federal Environmental Assessment process, Rainy River initiated the Provincial Environmental Assessment process through the submission of a draft Terms of Reference for public comment. A 30-day public comment period on the draft Terms of Reference was held between May 17, 2012 and June 16, 2012. A proposed Terms of Reference was issued for a 30-day public comment period from October 26 to November 26, 2012. This subsequent Proposed Terms of Reference or as amended, will set out the framework for the Provincial Environmental Assessment process once approved and is the first formal step in that Environmental Assessment process.

The Environmental Assessment document is currently in preparation based on the Feasibility Study designs and expected regulatory requirements, and is expected to be submitted as a draft for review during the second or third quarter of 2013. Rainy River is working closely with the Provincial and Federal governments to integrate the Environmental Assessment processes and regulatory schedules; however, there are challenges as there is no precedent in Ontario and the two (2) regulatory regimes are very different. The Ministry of the Environment and the Canadian Environmental Assessment Agency, as well as Rainy River are attempting to coordinate public consultation activities in order to minimize the effort required by stakeholders to be effectively engaged. These efforts are aimed at minimizing duplication and unnecessary delays. In December 2012, the Rainy River Gold Project was selected by the Federal/Provincial Regulatory Reform Working Group (RRWG) as one of a very limited number of projects across Canada, to receive enhanced alignment considerations and support from senior government officials. This senior support is expected to assist in maintaining the Project schedule.

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The Project environmental approvals schedule that has been incorporated into the overall Project schedule includes AMEC’s best estimate of the timeline needed to obtain environmental approvals based on published timelines, discussions with regulatory agencies, professional experience and precedents of other Ontario mining projects. Currently, all major mining projects in Ontario are subjected to the same coordinated Provincial/Federal Environmental Assessment process with the Provincial Individual Environmental Assessment process being a new approach to enhancing intergovernmental alignment in Ontario. While there is always a certain level of risk related to major project Environmental Assessment approval timing, the Rainy River Gold Project is subject to greater confidence level than there would have been prior to the promulgation of the Canadian Environmental Assessment Act, 2012. Environmental approvals to initiate construction must follow after and are dependent on the Environmental Assessment approvals.

The same body of information will be used to inform both the Provincial and Federal Environmental Assessment processes, culminating in a single Environmental Assessment report that meets both the Federal Environmental Impact Statement Guidelines and the approved Provincial Terms of Reference. After Rainy River issues the final Environmental Assessment report, the Federal and Provincial environmental assessment approval processes will continue in a parallel manner, to the extent possible, according to the regulated requirements.

Construction of the Rainy River Gold Project cannot proceed until the Federal and Provincial Environmental Assessments have been approved, and the appropriate regulatory approvals (as described below) have been attained for the Project component being constructed.

20.5.2.2

Environmental Approvals

A small number of Federal environmental approvals are anticipated to be required or are potentially required for the construction and operation of the Project as listed in Table 20-2.

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Table 20-2: Potential Federal Approvals


Approval/Licence	Agency Responsible	Description
Licence for an Explosives Factory (and Magazine) Explosives Act	Natural Resources Canada	Construction and operation of an explosives factory and magazine(s).

Authorization(s) for Harmful Alteration Disruption or Destruction of Fish Habitat Fisheries Act	Fisheries and Oceans Canada	Disruption to creeks and ponds supporting fish populations; approval for groundwater dewatering effects.

Schedule 2 Listing Metal Mining Effluent Regulation (MMER) Fisheries Act	Environment Canada	It is expected that the overprinting of waters frequented by fish by tailings and mine rock stockpiles (or other deleterious material) may be necessary and will also require a listing under Schedule 2 of the Federal MMER, pursuant to the Fisheries Act.

There are also five (5) primary Provincial agencies that will approve construction of the Rainy River Project: Ministry of Northern Development and Mines, Ministry of the Environment, Ministry of Natural Resources, Ministry of Transportation and the Ontario Energy Board.

The Ministry of Northern Development and Mines has a responsibility to ensure the orderly development of mineral resources in the Province, including responsibilities for mine closure activities. The Ministry of the Environment grants permits and approvals that address Project aspects related to water and air quality (including noise) and waste management. The Ministry of Natural Resource’s role is to ensure the protection and wise use of Crown resources (such as lakes and rivers, aggregates and Species at Risk) not otherwise disposed, such as through the Mining Act administered by Ministry of Northern Development and Mines. The Ontario Energy Board is an independent, quasi-judicial tribunal that is appointed by the Ontario government that has responsibility for energy-related approvals, including approval to construct transmission lines.

Table 20-3 summarizes the Provincial approvals anticipated to be required or likely to be required for the construction and operation of the Rainy River Gold Project. In some instances, multiple approvals may be needed of the same type.

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Table 20-3: Potential Provincial Approvals


Approval/Licence	Agency Responsible	Description
Environmental Compliance Approval – Air and Noise Environmental Protection Act	Ministry of the Environment	Approval to discharge air emissions and noise.

Environmental Compliance Approval – Industrial Sewage Works Ontario Water Resources Act	Ministry of the Environment	Approval to treat and discharge effluent such as for: mine/pit water, tailings management area.

Environmental Compliance Approval – Industrial Sewage Works Ontario Water Resources Act	Ministry of the Environment	Approval to treat and discharge effluent (such as for: sewage treatment, oil water separator).

Environmental Compliance Approval – Waste Disposal Site Environmental Protection Act	Ministry of the Environment	Operation of a waste transfer site.

Permit to Take Water Ontario Water Resources Act	Ministry of the Environment	Water taking from surface or ground water (open pit and other sources; multiple permits expected to be required).

Aggregate Permit Aggregate Resources Act	Ministry of Natural Resources	Establishment and operation of a sand and gravel pit/quarry, if applicable.

Approval to Commence Harvesting Operations (Forestry Resource Licence) Crown Forest Sustainability Act	Ministry of Natural Resources	Cutting of merchantable timber on Crown land.

Land Use Permit Public Lands Act	Ministry of Natural Resources	Land tenure for facilities constructed on Crown land (such as transmission line).

Species at Risk Screening Endangered Species Act	Ministry of Natural Resources	Management of activities related to Species at Risk.

Work Permit or other Approval Public Lands Act/Lakes and Rivers Improvement Act	Ministry of Natural Resources	Work/construction on Crown land, including below the high water mark of local watercourses and construction of dams.

Various Approvals (highway entrance, encroachment permit etc.) Highway Act	Ministry of Transport	Various engineering approvals, including those related to the relocation of Highway 600.

Closure Plan Mining Act	Ministry of Northern Development and Mines	For mine construction/production.

Leave to Construct Ontario Energy Board Act	Ontario Energy Board	Approval to construct a transmission line.

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20.6

Preliminary Environmental Impacts

The design of the Rainy River Gold Project is ongoing, as is the development of appropriate mitigation measures based on the feedback received during consultation to-date, and recognizing that design alterations may be needed as part of the regulatory process. Table 20-4 provides a preliminary summary of the potential environmental effects associated with the construction, operation and closure of the Rainy River Gold Project, by main project component. None of these environmental impacts are believed to be significant after application of the proposed mitigation measures.

Table 20-4: Preliminary Summary of Potential Environmental Effects


Undertaking Component	Potential Effect (Negative ‘-’; Positive ‘+’; D - direct; I - indirect; S – short term; L – long term)

Gold Mine	•	Reduction in localized air quality due to the release of particulate from mining activities and heavy equipment diesel emissions (-DS).

	•	Increase localized sound emissions as a result of intermittent blasting activities, heavy equipment operation and safety equipment (-DS).

	•	Alteration to the local terrain from excavation of the open pit, forming a permanent surface depression in the landscape (-DL).

	•	Potential for loss of aquatic habitat by the rerouting of West Creek to avoid the mine operation (-DL).

	•	Depression of the local groundwater aquifer by changes to the local landscape and mine dewatering activities (-DL).

	•	Potential effect on water quality in the Pinewood River from the release of treated effluent from the site, including treated mine water (-IS).

	•	Reduction in terrestrial habitat cause by the mine footprint development anticipated to be replaced by an open pit lake at closure (-DL).

	•	Temporary effect on local traffic by rerouting of Highway 600 to avoid the mine operation (-IS).

Buildings and Storage	•	Reduction in localized air quality and increase in localized sound emissions during construction (-DS).
Buildings and Storage	•	Reduction in localized air quality due to the release of emissions from the processing plant (-DS).

	•	Increase localized sound emissions as a result of processing plant and maintenance operations (-DS).

	•	Loss of local terrestrial habitat and/or quality of habitat including habitat for Species at Risk as a result of the process plant building and other buildings footprints (-DL).

	•	Potential effect on water quality in the Pinewood River from the release of treated effluent from the site, including treated process plant effluent and various wash water sources (-IL).

	•	Potential effect on the localized environment from accidents and malfunctions (-IS).

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Undertaking Component	Potential Effect (Negative ‘-’; Positive ‘+’; D - direct; I - indirect; S – short term; L – long term)

Stockpiles	•	Reduction in localized air quality due to the release of particulate matter from stockpiling activities and heavy equipment emissions (-DS).

	•	Reduction in localized air quality due to dust release from the stockpiles (-DS).

	•	Increase localized sound emissions as a result of heavy equipment operation, mineral waste deposition and safety equipment (-DS).

	•	Alteration to the local terrain through the forming of permanent stockpiles elevated about the existing landscape (-DL).

	•	Potential for loss of aquatic habitat by overprinting and/or rerouting of local creeks systems to accommodate stockpiling operations (-DL).

	•	Potential effect on water quality in the Pinewood River from the release of treated runoff and/or seepage from the stockpiles (-IL).

	•	Reduction in terrestrial habitat cause by the stockpile footprints (-DL).

Tailings Management Area	•	Reduction in localized air quality due to dust release from the tailings surface (-DL).
Tailings Management Area	•	Reduction in localized air quality due to the release of particulate matter from construction activities and heavy equipment operation (-DS).

	•	Increase localized sound emissions as a result of heavy equipment operation and safety equipment during Tailings Management Area dam construction (-DS).

	•	Alteration to the local terrain from the construction of a permanent facility raised above the surrounding landscape (-DL).

	•	Reduction in terrestrial habitat caused by the Tailings Management Area footprint (-DL).

	•	Potential for loss of aquatic habitat by local creeks and wetlands (-DL).

	•	Potential alteration of groundwater infiltration rates (-DL).

	•	Potential effect on water quality in the Pinewood River from the release of effluent and seepage from the Tailings Management Area (-DL).

On-site Infrastructure	•	Reduction in localized air quality and increase in localized sound emissions during construction (-DS).
On-site Infrastructure	•	Reduction in localized air quality due to dust release from roads and vehicle emissions (-DS).

	•	Loss and/or alteration of local terrestrial habitat and/or quality of habitat including for Species at Risk, as a result of the infrastructure footprints (-DS/-DL).

	•	Potential effect on the localized environment from accidents and malfunctions (-IS).

Realignment of existing Highway 600	•	Reduction in localized air quality and increase in localized sound emissions during construction (-DS).
	•	Alteration to the local terrain (-DL).
	•	Potential for minor loss of aquatic habitat from culver/bridge installation (-DL).

	•	Reduction in terrestrial habitat and/or quality of habitat cause by the altered highway footprint (-DL).

	•	Temporary inconvenience to local landowners during construction/rerouting activities (-DS).

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Undertaking Component	Potential Effect (Negative ‘-’; Positive ‘+’; D - direct; I - indirect; S – short term; L – long term)

Overall RRGP	•	Direct local economic benefits, employment and business opportunities, direct expenditures and taxes (+DS/+DL).

	•	Indirect local economic benefits, spin-off employment and business opportunities; spin-off expenditures and taxes (+IS/+IL).

	•	Direct regional economic benefits, employment and business opportunities, direct expenditures and taxes (+DS/+DL).

	•	Indirect regional economic benefits, spin-off employment and business opportunities; spin-off expenditures and taxes (+IS/+IL).

	•	Direct Provincial economic benefits, employment and business opportunities, direct expenditures, taxes and royalties (+DS/+DL).

	•	Indirect Provincial economic benefits, spin-off employment and business opportunities; spin-off expenditures and taxes (+IS/+IL).

	•	Direct Federal economic benefits, employment and business opportunities, direct expenditures and taxes (+DS/+DL).

	•	Indirect Federal economic benefits, spin-off employment and business opportunities; spin-off expenditures and taxes (+IS/+IL).

	•	Extra demand on existing infrastructure and social services in the region (-IS).

*	For the purposes of this table, short term has been defined to include the construction, operation and active closure phase of the RRGP.

20.7

Preliminary Reclamation Plan

Reclamation will be carried out in accordance with all applicable regulations. The primary closure considerations are to protect human health and the environment, and provide enhancement to the affected environment. This includes appropriate long-term management of potentially acid generating/metal leaching materials and any affected waters. The objective of final reclamation for the Rainy River Gold Project is to return the site to a productive condition on completion of mining activities. It is fully expected that site aesthetics and land use will need to be considered acceptable in regards to government regulations and industry standards as well as Aboriginal and public expectations subject to practical considerations. These needs have attempted to be anticipated in the proposed preliminary reclamation plan and associated costing.

The Ontario Mining Act requires that a closure plan be filed for any mining project before the construction of the Project is initiated. It also requires that financial assurance to the Province to provide sufficient funds to support the activities required by the closure plan be submitted at the same time as the closure plan.

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There is the potential that some industrial facilities could remain on the site after closure for other industry use, but complete reclamation has been assumed for the purposes of this Feasibility Study, including removal of the 230 kV transmission line connection. The relocated Highway 600 is a permanent change to the regional road network and will remain in place after operations cease.

Table 20-5 summarizes a preliminary closure approach for the Rainy River Gold Project. As much as possible, work will be completed progressively during operations, which is considered an industry best management practice.

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Table 20-5: Preliminary Reclamation Approach


Project Element		Preliminary Approach

Open Pit	•	Remove all infrastructure and equipment within the open pit.

	•	Shape and revegetate overburden slopes to a stable condition to increase physical stability and provide natural organic matter to the pit lake in the longer term.

	•	Allow pit to fill naturally by means of seepage and runoff inputs from the local area including the mine rock stockpiles, together with possible supplemental pumping from the Pinewood River.

	•	Provide fencing as an interim measure for public safety until pit no longer poses a safety hazard.

	•	Block the entrance to the open pit using boulder fencing or other means.

	•	West Creek will likely be returned to its original path and will be routed through the ultimate pit lake.

Underground Mine	•	Remove all infrastructure and equipment with value within the underground workings.

	•	Allow workings to flood naturally via groundwater seepage.

	•	Block the entrances to underground (capping vertical openings and blocking ramp by overfilling with boulders).

Buildings, Machinery, Equipment and Infrastructure	•	Salvageable machinery, equipment and other materials will be dismantled and taken off site (sale or reuse).
	•	Remaining items will be cleaned of oil and grease and deposited within an approved landfill.
	•	Equipment containing hydrocarbons that cannot be readily cleaned will be trucked off site for recycling or disposal at a licensed facility.

	•	Above grade concrete structures (including abandoned structures, unless designated as heritage features) will be broken and reduced to near grade with rebar and will be cut flush with the surface.

	•	Concrete structures will be infilled with clean mine rock, if needed.

Roads	•	Some site roads will be scarified, edges resloped as appropriate, covered with overburden and seeded when no longer needed. Others may be kept for access and/or returned to the Municipality as part of zoning considerations.

Pipelines and On-site Transmission Lines	•	Pipelines (such as tailings, freshwater and effluent) will be purged, dismantled, and disposed of at an approved landfill.

	•	On-site power lines and associated materials without value will be dismantled and deposited in an approved landfill.

	•	Other power equipment and materials including oil-filled transformers will be taken off-site for sale or reuse.

	•	The 230 kV transmission line will be dismantled, recycled if possible, or otherwise deposited in the landfill if no future use is foreseen (and as currently anticipated).

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Project Element		Preliminary Approach

Petroleum Products, Chemicals and Explosives	•	All petroleum products and chemicals remaining will be removed from the site and transported to a licensed facility for disposal.

Contaminated Soil	•	Soil found to exceed acceptable criteria will be bioremediated on site or handled otherwise according to regulatory requirements.

Ponds and Other Water Structures	•	Water structures not needed for longer term water management will be removed, if necessary, to ensure proper site drainage.

Tailings Management Area	•	Tailings to be maintained in partial permanent water cover condition with soil cover on the margins.

	•	Overflow spillways within the dams will be deepened, if needed.

	•	Perimeter and discharge ditches will be left in place and protected from erosion, where needed.

Overburden/Non- Potentially Acid Generating Mine Rock Stockpile	•	Stockpiled overburden will be used for reclamation on site, as needed, including as a cap over the potentially acid generating mine rock.
	•	Remaining stockpile will be resloped and revegated once no longer required.

	•	Non-acid generating rock may be covered with overburden and revegetated to enhance aesthetics.

Encapsulated Potentially Acid Generating Mine Rock Stockpile	•	Stockpile will be encapsulated within a multi-layer cover prescibed for both stockpile tops and sideslopes:
		• Fine grained compacted overburden layer directly on the mine rock;
		• Layer of non-potentially acid generating rock;
		• Final fine-grained compacted silt-clay overburden layer.

	•	Revegetation of the overburden cap.

	•	Runoff and seepage from the stockpile will be collected, and if necessary, discharged into the open pit.

Once a decision has been made to permanently close the site, it is anticipated that the major closure activities would be completed within a period of approximately two (2) years, if not already completed progressively.

Monitoring of various site aspects such as water quality, revegetation success and Tailings Management Area dam stability is expected to continue over an extended period of time. Monitoring and management of runoff from the waste rock stockpile will also be conducted to ensure that site drainage is compatible with downstream receiving waters.

Closure costs have been prepared based on this approach and are included in the sustaining capital and financial analysis sections of this study (Sections 21 and 22, respectively).

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21.

CAPITAL AND OPERATING COSTS

The capital and operating costs developed in this Study are based on the construction of a Greenfield open pit/underground mine and process plant facility having a nominal daily treatment capacity of 21,000 tpd of mineralized feed. The capital and operating cost estimates related to the open pit mine, concentrator and site infrastructure have been developed by BBA and Merit Consultants. Costs related to the underground mining operation have been estimated by Golder Associates (Golder). AMEC provided tailings and water management material quantities, the water management plan and site closure costs. BBA and Merit consultants consolidated the cost information from all sources to determine the overall Project capital cost.

Type and Class of Estimate

The overall capital and operating costs developed in this Study meets the AACE Class 3 requirements and has an accuracy range between -10% and +15%. The Capital Cost Estimate of this Feasibility Study forms the basis for overall project budget authorization and funding. It is the “Control Estimate” against which subsequent phases of the Project will be compared.

The Capital Cost Estimate abides by the following criteria:

•

Assumes contracts will be awarded to reputable contractors on a lump sum basis and an open shop environment;

•

Pre-production capital costs are expressed in constant Q4 2012 Canadian dollars (“CAD”), with an exchange rate at par with the US Dollar (“USD”);

•

Provides a common basis for classifying all types of facilities and processes;

•

Defines the major characteristics used to classify cost estimates;

•

Primarily classifies cost estimates based on a measurable degree of engineering completion; and

•

Reflects general accepted practices in the cost engineering profession.

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21.1

Capital Cost Summary

The projected pre-production (initial) capital cost and contingency for the Project is estimated to be $713.3M. Open pit sustaining capital costs are estimated to be $321.9M. Underground development phase capital is estimated at $67.8M with underground sustaining capital estimated at $94.6M. Taxes and land acquisition, permitting, licensing and project financing costs are not included in the cost estimate. The Project capital cost summary is outlined in Table 21-1. The capital cost breakdown descriptions are outlined in the following sections.

Table 21-1: Project Capital Cost Summary


Area Description	Pre-Production Capital Costs ($M)		Sustaining Capital Costs ($M)
Overhead Power Line		11.1		—
Highway 600 Replacement		10.1		—
Open Pit Pre-Stripping (Overburden)		43.2		70.7
Open Pit Mine Equipment		26.2		161.5
Open Pit Rock Excavation¹		45.1		62.7
Underground – Development Phase Capital²		—		67.8
Underground – Sustaining Capital³		—		94.6
General Site Infrastructure		81.2		6.0
Process Facilities		283.7		1.5
Tailings and Water Management		44.9		23.6
Housing and Equipment Salvage Value		—		(64.2	)
Reclamation and Closure Costs		—		60.1
Indirect Costs		112.8

Subtotal		658.3		484.3

Contingency		55.0

Total Capital		713.3		484.3

¹	Capitalization of waste excavated rock and ore during the pre-production period and operating years with stripping ratios higher than 3.1 (LOM).

²	Funded through internal cash flosws, this is the capital required in the development phase of the underground mine, consisting of equipment and infrastructure, as well as vertical and horizontal development.

³	Funded through internal cash flows, this is the sustaining capital required for the underground mine, consisting of equipment and infrastructure, as well as vertical and horizontal development.

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21.2

Overhead Power Line

The cost of the high voltage power line from the local power grid was estimated from budgetary quotations reviewed by SanZoe Consulting Inc.

21.3

Highway 600 Replacement

TBT Engineering provided engineering input for the road design to meet MOT (Ministry of Transportation) requirements. The cost of Highway 600 realignment was estimated by Merit based on bulk earthwork quantities provided by BBA.

21.4

Open Pit Mining Capital Cost Estimate

The open pit pre-production and sustaining capital costs include overburden pre-stripping, mine equipment and ore and waste production costs occurring during the pre-production period or above the LOM stripping ratio of 3.1:1 (waste : ore). The open pit pre-production and sustaining capital costs for pre-stripping, waste removal and equipment estimated by BBA total $114.5M and $294.9M, respectively.

Open Pit Mining Equipment Costs

Open pit mining equipment has a pre-production capital cost of $26.2M and $161.5M in sustaining capital costs. The total cost for mine mobile equipment, aggregate plant and mobile equipment dispatch system, including interest is $187.7M. The equipment quantities and cost without interest are presented in Table 21-2.

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Table 21-2: Open Pit Mine Equipment Capital Cost


Equipment	Type / Capacity	Quantity	Unit Price		Total Cost
			(‘000$)		(‘000$)
Mine Fleet
Haul Truck	226 t	19		4,229		80,359
Diesel Shovel	28 m³	1		12,300		12,300
Electric Shovel	28 m³	2		12,100		24,200
Blast Hole Drill	8.5”	3		2,646		7,937
Support Fleet
Large Wheel Loader	15 m³	1		5,381		5,381
Motor Grader	300 HP	2		1,122		2,244
Hydraulic Excavator	523 HP	1		1,260		1,260
Track Dozer	580 HP	7		1,706		11,944
Vibratory Soil Compactor	156 HP	1		189		189
Small Wheel Loader (Used)	6.4 m³	1		300		300
Auxiliary Fleet
Aggregate Plant		1		1,000		1,000
Boom Truck	19 t 70’ boom	2		154		308
Dewatering Pump	100 HP	2		82		164
Fuel/Lube Truck	19,000L fuel	2		1,175		2,350
Hydraulic Crane, truck-mounted	150 t	1		800		800
Lighting tower 4-post of 1000 w / diesel generator	4-1000 w	6		10		60
Mini Bus	12-seater	2		48		96
Mobile Pump	150 HP	4		85		341
Pickup Truck Crew Cab		12		51		621
RC Drill	4.5”	2		950		1,900
Scraper	400 HP	2		400		800
Service Truck	250 HP	2		130		260
Stemming Loader	2.7 m³	1		200		200
Tire Changer, truck-mounted		1		233		233
Tow Truck Unit	1,025 HP	1		2,560		2,560
Water Truck	30,000 L	1		876		876
Water Truck (Used)	60,000 L	1		1,870		1,870
Small Fuel / Lube Truck	16,000 L fuel	1		325		325
Other Equipment
Geotech Monitoring Equipment /Survey Equipment		1		200		200
Dispatch System		1		457		457

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Mining equipment quantities and costs have been developed based on the mine plan. Mining equipment costs are based on quotes requested during this Study and from BBA’s recently updated database of supplier pricing. In order to reduce initial mining equipment costs, it is assumed that Rainy River Resources will finance the equipment purchases with the equipment suppliers.

The capital cost for the open pit mining equipment is calculated on the basis that a 9% down payment will be placed on each piece of equipment purchased and financed over an 8-year period at an interest rate of 4.5%. Down payments and financing cost incurred during pre-production Years -2 and -1 are included in the pre-production capital cost; however, the completion of pre-production equipment financing is considered to be a sustaining capital cost. All mining equipment required during the life of operation is considered to be sustaining capital.

Sustaining capital required to support mining operations also include costs associated with the phased construction of ditches and settling basins around the perimeter of the waste piles that are required for compliance to regulations concerning total suspended solids effluent water quality.

For this Study, it is assumed that the installations for servicing mining equipment will be built during the pre-production period. These installations will comprise a permanent truck wash station with mud settling basins and a permanent truck shop building for mine equipment maintenance consisting of a six-bay garage.

21.5

Underground Mining Capital Costs

The total capital and operating costs were estimated from first principles using a cost model built by Golder Associates for this Project. All currency is in Canadian dollars. Quotes were obtained for all major equipment and consumables. Contingencies were calculated based on the type of cost (multiple quotes, single quote, or factored). Overall, a 10% contingency on mine development drifting and fixed equipment installation cost is used for this Study. This was discussed with Rainy River, and it was decided that a 10% contingency, based on development meterage, would be applied to only the mine development capital cost.

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Mine Capital Costs

The mine capital costs consist of mobile and stationary equipment and mine development. Multiple quotes were obtained from equipment manufacturers for the major mobile equipment; single quotes were obtained for the rest. Construction estimates were made for the ventilation and dewatering infrastructure.

Mobile Equipment

The quantity, unit cost, and total cost of the mobile equipment required for the Rainy River Underground is shown in Table 21-3. The capital costs of the development and production equipment were estimated using a leasing formula provided by Rainy River over a period of eight (8) years, which matches the useable equipment life. This leasing arrangement has improved the net present value of the underground mine, but has increased the capital cost of the leased equipment by an average of 6%.

The capital cost of the support equipment was included as a purchase. It is estimated that the RRU will require a total of $31.4M of equipment over the life of the mine.

Table 21-3: Mobile Equipment Capital Cost Estimate


Equipment	Type	Quantity	Unit Cost (’000$)		Total Cost (’000$)
Development Equipment
Development Jumbo	2-boom	2		1,100		2,200
Development LHD	4 m³	2		980		1,960
Development Haul Truck	15 m³	3		910		2,730
Development Bolter		3		855		2,565
Development Anfo Charger		2		485		970
CAF Production
Production Jumbo	2-boom	1		1,100		1,100
Production Truck	27 m³	1		1,545		1,545
Backfill LHD	4 m³	1		980		980
Production LHD	5 m³	1		1,370		1,370
Production Anfo Charger		1		490		490
Production Bolter		1		860		860

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Equipment	Type	Quantity	Unit Cost (’000$)		Total Cost (’000$)
Longhole Production
Production LHD	5 m³	2		1,370		2,740
Production Truck	27 m³	2		1,545		3,090
Surface Loader	8.4 m³	1		1,360		1,360
Emulsion Loader		1		500		500
Longhole Drill Rig		2		1,110		2,220
Backfill LHD	4 m³	1		980		980
Support
Scissor Lift		2		360		720
Boom Truck		2		320		640
Grader		2		235		470
Personnel Carrier		6		105		630
Slurry Plant		3		410		1,230
Lube Truck		1		105		105

TOTAL						31,455

Equipment purchases were scheduled to account for the planned equipment life and the potential to use equipment in different mine areas. For example, a bolter can be used in either a CAF production stope or a development heading. The underground development at Rainy River is scheduled over years, with seven (7) years of steady use (continual development) and four (4) years of intermittent use (e.g., attack ramp development). Cut and fill mining starts approximately two (2) years after the company development and extends for eight (8) years, overlapping with final years of intermittent development. It is felt that the combined CAF and development equipment will be able to maintain the combined CAF and development schedules even if the development equipment has been scheduled beyond its estimated life.

Production from the longhole stopes at the Rainy River is scheduled for 10 years, including a 3-year ramp up and a 1-year ramp down period. Productivity estimates indicate that only one (1) longhole drill and one (1) LHD are required to meet production targets. To reduce the impact of unexpected downtime of the remote LHDs and the longhole drills on daily production, it was decided to include the cost of an additional longhole drill and an addition remote-LHD in the capital costs.

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Three (3) trucks are required to meet production targets from the Rainy River underground. The purchase of these trucks has been scheduled to match the ramp up in production from the underground, which means that the last year of production from the longhole area will be in the ninth year of the newest truck (45,000 hours). It is felt that the planned truck fleet will be sufficient considering the reduced production requirements (approximately 600 tpd) in the final year of the mine and additional purchases will not be required.

Stationary and Auxiliary Equipment

Table 21-4 shows the estimated cost of stationary equipment for the RRU. Included in these costs are items such as the sumps, ventilation fans and infrastructure, additional ground support, infrastructure for the portals, and mine rescue gear. An allowance has been made to equip and prepare the small underground shop and warehouse facilities, and for ancillary gear such as survey, safety, office, and laboratory equipment.

Table 21-4: Estimate of the Stationary and Auxiliary Equipment Capital Cost


Equipment	Total Cost (’000$)
Main Sump		3,300
17 East Sump		2,800
433 Sump		2,800
ODM Sump		2,800
Air Compressor		150
Ventilation Fans		1,010
Ventilation Infrastructure		320
Heaters		2,100
Refuge Stations		300
Electrical Distribution		5,000
Portals		1,000
Mine Rescue Equipment		150
Cap Lamps		120
Warehouse/Shop		1,000
Ancillary Equipment		500

Total		23,350

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Equipment Operating Costs

The equipment operating costs for major equipment were obtained from supplier quotes and for minor equipment from InfoMine’s Mine and Mill Equipment Cost Calculator, Costmine (InfoMine, internet site). They are shown in Table 21-5 , and include wear parts, lube, and tires. Fuel consumption is not included in these costs.

Table 21-5: Equipment Operating Cost used at the Rainy River Underground


Equipment		Operating Cost ($/h)
5.6 m³LHD		90
27 m³Truck		62
4 m³LHD		102
15 m³Truck		42
Jumbo		31
Bolter		40
Longhole Drill		45

In discussion with Rainy River Resources, it was decided to use a 40,000 hour (8 years) useable equipment life for the major mobile equipment as opposed to the conventional 25,000 hours (5 years). The variable parts cost at various engine hour intervals for the major mobile equipment was quoted by a supplier. These costs include three (3) rebuilds and are factored based on Golder’s experience.

Underground Mine Workforce

The workforce was estimated based on mine activities and mobile equipment quantities, industry standard ratios for maintenance personnel, and input from Rainy River. The workforce is separated into staff and hourly labour, and the yearly base rate, markup, and yearly cost are shown in Table 21-6 and Table 21-7 for both types. The markup includes the benefits and burdens and was chosen to be consistent with the rest of the Feasibility Study. The hourly labour costs were calculated based on a 2,100-hour work year and were benchmarked against operating underground gold mines in the region.

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Table 21-6: Peak Staff Complement and Cost for the Rainy River Underground


Staff	Peak Quantity (2022)	Yearly Base Rate ($)		Markup (%)	Yearly Annual Cost ($)
Development Superintendent	1		120,000	30		156,000
Production Superintendent	1		120,000	30		156,000
Mine Captain	1		120,000	30		156,000
Mine Supervisor	8		110,000	30		1,144,000
Chief Engineer/Geologist	2		120,000	30		312,00
Mine / Geotechnical / Ventilation Engineer / Geologist	5		100,000	30		650,000
Engineering / Geology Technician	4		65,000	30		338,000
Health and Safety Officer	1		65,000	30		84,500
Clerk	1		40,000	30		52,000
Maintenance Planner	1		120,000	30		156,000
Shop Supervisor	1		110,000	30		143,000

Total	26					3,347,500

Table 21-7: Peak Hourly Complement and Cost for the Rainy River Underground


Labour	Peak Quantity (2022)	Yearly Base Rate ($)		Markup (%)		Yearly Annual Cost ($)
Miner	24		76,650		30		2,391,000
Operator	88		70,080		30		8,017,000
Labourer	16		65,700		30		1,367,000
Mechanic	12		70,080		30		1,093,000
Electrician	12		70,080		30		1,093,000
Technician	1		41,610		30		54,000

Total	153						14,015,000

Mine personnel were assigned to either capital or operating cost depending on their task. For example, the cost of a miner developing the main ramp will contribute to the mine capital costs, whereas the cost of a miner developing a stope overcut will contribute to the mine operating costs.

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The staff at the RRU will be working a 40-hour, 5-day work week, and the hourly workforce will be working 11 hours per day, on a 2-on / 2-off schedule. Initially, a small complement of staff will be on site to prepare the mine plans and supervise the contractor. As the mine deepens, the staff level increases with additional technical personnel. Once the mine starts development, additional staff and hourly employees are hired to begin development, maintenance and production activities. At peak production, Year 2022, there will be 26 staff and 153 hourly employees.

Consumables

The unit costs of the main consumables are shown in Table 21-8. The cost of explosives, propane and cement were determined through supplier quotes. The costs of power and diesel fuel were provided by Rainy River. Additional consumables (not listed) including ground support, additional blasting supplies, drilling wear parts, pipes, and ventilation ducting were obtained from quotes or Golder’s database of costs.

Table 21-8: Cost of Major Consumable Items


Item	Cost		Unit
Power		0.06	$/kwh
Diesel		0.85	$/l
Ammonium Nitrate/Fuel Oil		0.85	$/kg
Emulsion		1.05	$/kg
Propane		0.90	$/l
Cement		260	$/tonne

Lateral Development

The majority of the RRU will be developed with company crews.

Table 21-9 shows the average cost per metre for the various headings designed. The total development cost is shared between capital and operating costs and approximately one-third of the development cost is operating and two-thirds is capital, based on length. The unit cost includes machine operating costs, weighted average labour and consumables. The contingency was estimated using the consumable cost for a 5 m wide by 5 m high drift.

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Table 21-9: Unit Company Development Costs


Description (width x height)	Unit Cost ($/m)
4.5 m x 4.5 m drift		4,300
5 m x 4.5 m drift		4,000
3 m x 5 m drift		3,700
4 m x 4 m drift		3,800
6.5 m x 5 m drift		4,400
4.5 m x 3.5 m drift		3,700
5 m x 5 m drift - contingency		1,300

Contractors

Contractors are scheduled to develop the main ramp and excavate all capital vertical excavations. Quotations for lateral development were obtained in 2010 and, based on recommendations from the contractor, were factored to 2012 rates. Raise boring quotes were obtained for the size of raise anticipated for the RRU. The contractor costs are summarized in Table 21-10 and include labour and machine time. Fuel, power, ground support and, in the case of the raise bores, waste removal, are considered a company cost and are not included here.

Table 21-10: Contractor Development Cost


Description	Contractor Cost ($/m)		Owner’s Cost ($/m)		Total Cost ($/m)
5.5 m x 6 m drift - contracted		6,700		800		7,500
Raise (2 m diameter) (waste)		3,800		600		4,400
Raise (2.5 m diameter) (waste)		3,900		600		4,500
Raise (3 m diameter) (waste)		4,550		700		5,250
Raise (4 m diameter) (waste)		4,600		700		5,300
Raise (4.5 m diameter) (waste)		7,300		1,000		8,300

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Underground Capital Cost Summary

Table 21-11 shows a summary of the capital cost over the life of the RRU. The contractor cost includes both lateral and vertical development, the latter of which is required until the 433 area is completely developed in year 2024. The equipment costs are paid until the final year of production as the equipment is leased. Year 2021 has very little capital cost as the majority of the development occurs in the ODM Zone preparing the stopes for diamond drilling. Overall, it is estimated that the RRU will require a capital investment of $162M.

Table 21-11: Rainy River Underground Capital Cost Summary


Year	Development¹ (‘000$)		Development Equipment (‘000$)		Production Equipment (‘000$)		Support Equipment (‘000$)		Total (‘000$)
2016		8,500		0		0		1,100		9,600
2017		24,400		1,080		0		9,890		35,370
2018		18,780		1,420		1,250		1,370		22,820
2019		15,280		2,330		1,710		280		19,600
2020		15,790		1,450		1,840		2,940		22,020
2021		750		1,690		2,140		2,130		6,710
2022		8,780		1,560		2,230		2,840		15,410
2023		8,930		1,350		2,230		0		12,510
2024		6,200		1,760		2,230		2,800		12,990
2025		50		900		2,050		0		3,000
2026		70		680		520		0		1,270
2027		20		0		520		0		540
2028		10		0		520		0		530

Total		107,560		14,220		17,240		23,350		162,370

¹	“Development” here refers to both development phase and production phase vertical and lateral development.

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21.6

Process Plant and Site Infrastructure Capital Costs

The process plant, site infrastructure and tailings management area (“TMA”) pre-production capital cost is estimated to be $598.8M, including Direct Costs, Indirect Costs and contingency. The costs are detailed in Table 21-12.

Table 21-12: Infrastructure, Process Plant and TMA Pre-Production Costs


Area	Total Cost ($M)
Direct Costs
Infrastructure		102.4
Process Plant		283.7
Tailings Management Area		44.9

Subtotal		431.0

Indirect Costs

Subtotal		112.8

Contingency

Subtotal		55.0

TOTAL		598.8

21.7

Process Plant Capital Costs

The design of the crusher area, the crushed ore stockpile area and the concentrator area has largely been based on BBA’s experience on recent projects. The site plan and General Arrangement (GA) drawings developed in this Study have been used to estimate quantities and generate Material Take-Offs (MTOs) for all commodities. Equipment costs have been estimated using budgetary proposals obtained from suppliers for most process equipment. For process and mechanical equipment packages, equipment datasheets and summary specifications were prepared and budget pricing obtained from suppliers. For packages of low monetary value, pricing was obtained from BBA’s recent project data, when available. A detailed equipment list was developed with equipment sizes, capacities, motor power, etc. Related infrastructure has been estimated by BBA based on the site plan developed.

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21.8

Tailings and Water Management Capital Costs

The main components of sustaining capital related to the TMA and water management include:

•

Phased construction of TMA dams based on the tailings management strategy developed by AMEC;

•

The construction of an artificial wetland; and

•

An additional tailings pump booster station and tailings pipeline.

During life of the operation, an estimated sustaining capital of $23.6M is required for necessary additions or improvements to the water management and tailings management area.

21.9

Rehabilitation and Mine Closure Costs

Progressive rehabilitation and mine closure costs have been estimated by AMEC to be $35.5M and $24.6M, respectively. Regulatory guidelines require these costs be considered as a Financial Assurance. This cost is taken into account in the Financial Analysis but is not considered part of the pre-production capital cost.

21.10

Direct Cost - Basis of Estimate

Mining

Mining direct costs include the mining fleet equipment, as discussed in Section 21.4 and 21.5., the dispatch system, geotechnical monitoring and survey equipment, and the costs associated with the open pit mine development (earth and rock excavation).

Civil

Earthwork quantities were estimated from drawings, topographical data and geotechnical information.

Concrete

Preliminary design sketches were used to develop the concrete and embedded steel quantities. Unit rates, including formwork and rebar, were estimated from similar projects overseen by Merit Consultants.

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Structural

Material was priced from current steel market values and benchmarked against current projects.

Architectural

Siding and roofing quantities were estimated from General Arrangement drawings. Pricing based on Merit Consultants’ references on recent data from similar projects.

Mechanical

A platework list was developed with sizing, weights and surface areas including lining requirements. An HVAC equipment list was developed and fire protection and HVAC material take-offs (MTOs) were taken from layout and elevation drawings.

Piping

Complete piping diagrams were prepared. Pipe lining requirements were also categorized. Lengths for each line shown on the piping diagrams were determined from layout drawings. Material pricing for carbon steel and rubber-lined piping was obtained from supplier proposals.

Electrical

An equipment list including capacities and sizing was subsequently developed from the single line diagrams. MTOs for electrical bulk quantities were derived from cable schedules and runs, including cable trays routing layouts. Datasheets and were prepared for all major electrical equipment and components, and budget pricing obtained from suppliers. For electrical equipment of lower value, BBA’s historical cost data was used.

Automation/Telecommunications

A detailed instrumentation list was developed from the process flow diagrams.

Construction Labour Rates

All estimated costs for labour are based on ten (10) hours per day, seven (7) days per week, for a total of seventy (70) hours worked per week. Employee rotations of 3 (three) weeks of work and 1 (one) week of rest is expected. There is no allowance for evening shifts except for the mine pre-development which will be executed by Rainy River Resources mining crews.

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Separate methodologies for bulk earthwork and other trades were used to determine the hourly construction crew rates used in this estimate.

For bulk earthworks, budgetary quotations were obtained from five (5) qualified contractors. These quotes were developed by the contractors from material take-offs (MTOs) generated during the PEA Update Study. After a commercial analysis was conducted, a quote was selected as the pricing basis for the estimate. For every work area and work element, the crew rate including cost of operation of heavy machinery was calculated based on the selected quote.

For the other trades, two (2) general contractors were contacted to develop a single combined crew rate, averaging the cost of all trade work on the Project, based on the assumption that all hourly workers are hired by open shop contractors.

This average crew rate is built from four (4) components as follows:

1.	The direct cost is calculated from the rates of direct labour that are labourers, apprentices, journeymen and lead-hands, foremen, general foremen, and superintendents;

2.	The indirect cost is calculated from the rates of supervisors, i.e., the contractor’s project manager, engineers, quality control, planners, secretarial;

3.	Living Out Allowance estimated at $120 per day + markup; and

4.	Contractor construction equipment costs estimated at 15% of total crew rate.

The direct and indirect cost components are calculated on an assumption of 70 hours per week considering 40 hours at regular rate and 30 hours applying an overtime multiplier of 1.5 to the regular rate. General foremen and superintendent rates are calculated based on a 73.5 hours per week basis with 40 hours at regular rate and 33.5 hours applying an overtime multiplier of 1.5 the regular rate. Both the direct and indirect costs components include provisions for small tools, insurance, overhead and profit and travel turnaround costs. The contractor’s direct personnel are estimated to use 73% and contractor’s indirect personnel 27% of the construction effort.

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Productivity

Labour productivity is often the greatest risk factor, and uncertainty in cost and scheduling. The two (2) most important measures of labour productivity are:

1.	The effectiveness with which labour is used in the construction process; and

2.	The relative efficiency of labour, doing what it is required, at a given time and place.

Important factors affecting productivity on a construction site (such as site location, labour turnover, health and safety consideration, weather conditions, supervision, etc.) were considered to calculate the labour productivity factors shown in Table 21-13.

Table 21-13: Labour Productivity Factors


Calculated Productivity Loss Factor
Trade	Factor
Electrical		1.606
Automation/Telecom		1.593

The productivity factors are applied on installation costs evaluated from first principles.

Costs for trades not identified in Table 21-13 were evaluated by obtaining quotes from specialized contractors aware of anticipated site conditions and not from first principles. Applying a calculated productivity loss ratio was therefore not required.

Winter conditions expected between December 1 to March 31, are taken into consideration within the aforementioned productivity factors and are also considered in the first year for civil, concrete and steel works, as indicated in the Project Execution Schedule presented in Section 24 of this Study.

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21.11

Indirect Cost - Basis of Estimate

Indirect Costs were estimated jointly by BBA and Rainy River Resources and are divided into several categories as shown in Table 21-14.

Table 21-14: Indirect Costs


Indirect Costs	Millions ($ M)
EPCM		44.0
Owners’ Cost		18.2
Construction Operations Cost		6.8
Third Parties		4.4
POV and Commissioning		4.9
Construction & Commissioning Spares – Freight – Supplier Representatives		17.8
Capital Spares		8.7
Construction Infrastructure		8.0
Contingency		55.0

TOTAL INDIRECT COSTS		167.8

EPCM

EPCM Service Costs were developed based on BBA’s reference data for projects of similar size and schedule.

Owners’ Costs

Rainy River Resources prepared an itemized list with budget allowances for the Owners’ costs. Owner costs include items such as, but not limited to: administration and management personnel, project management team personnel, housing, construction insurance, purchasing costs, environmental expenses and training. Personnel hired during the pre-production period were also included in owners’ costs.

Construction Operations Cost

Construction operation costs include the construction and maintenance of temporary worker facilities required during the project construction period. An itemized list with budget allowances

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was developed by BBA and includes the costs for site distribution of temporary construction power, access roads to the temporary construction facilities, a telecommunication tower and other related equipment.

Supplier representation during construction is estimated based on projects of similar magnitude.

Third Parties

Cost of sub-consultants and other third parties were estimated based on projects of similar size.

Pre-Operational Verifications (POV) and Commissioning

POV & commissioning costs were estimated based on projects of similar size and include specialist professional services required to assist Rainy River Resources in the final acceptance of the work and plant start-up as well as an allocation of construction manpower resources for miscellaneous commissioning adjustments.

Construction and Commissioning Spares, Freight

Costs related to construction and commissioning spare parts were estimated to be 4% of the equipment purchase costs. A freight forwarder provided a quote for equipment freight costs. The quote was approximately equivalent to 6.75% of the equipment cost.

Construction Infrastructure Costs

Costs for mobile equipment and vehicles used during construction were estimated based on projects of similar size.

Capital Spares

BBA prepared an itemized list with budget allowances for the Capital Spares.

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Contingency

Contingency provides an allowance for undeveloped details within the scope of work. It does not account for labour disruptions, weather-related impediments, changes to the scope of the Project, price escalation nor currency fluctuations.

In order to establish and adequate contingency estimate, BBA conducted a probabilistic risk analysis on the capital cost estimate. Each estimate item was categorized in terms of precision and source of cost information.

The quality and nature of how costs were established for the materials and equipment portions of the estimate have a level of meticulousness that instills confidence. In regards to installation, productivities were based on experience, with some degree of input from contractors to provide additional comfort. The quantities provided by the engineers are in the expected range for a plant of this size and capacity.

Given a quantitative risk input for each item, a simulation software performing a monte-carlo probabilistic analysis yielded a data set showing the likeliness of achieving the Project budget under a given total, shown in Figure 21-1.

LOGO

Figure 21-1 - Cumulative Probability

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Figure 21-1 indicates a 95% chance of achieving the project under the estimated pre-production capital cost of $713M. This motivated the selection of $55M as the required contingency allocation.

Near the end of the Feasibility Study, an exercise with a different approach was performed by Revay, an independent reviewer hired by Rainy River Resources. Revay’s team performed a stochastic analysis (Monte Carlo) based on epistemic and aleatoric risks. The former is based on cost estimate uncertainty ranges as determined in sessions with personnel from BBA, AMEC and Merit. The latter is the cost of event risks. The event risks were identified during a risk review workshop animated by Revay where personnel involved in the Study as well as independent auditors participated.

The total risk cost estimated by Revay to meet a class 3 estimate with a precision level of ± 15% is $41.6M. Revay’s risk evaluation, once added to the base capital cost estimate, yielded a total project initial capital cost of approximately $710M which is within 1% of BBA’s estimate. Based on Revay’s evaluation, the $55M contingency allocation was maintained.

21.12

Process Plant and Infrastructure Sustaining Costs

Sustaining capital for the process plant is approximately $1.5M and includes allowances for mobile equipment, shop equipment and provisions for three (3) electrowinning cells to be added in Year 2, to account for higher than average silver grades in subsequent years.

Approximately $23.5M in infrastructure sustaining capital is allocated to account for the dam construction and the tailings line expansion in Year 4. A provision of $6M for project fencing is anticipated over the first two (2) years of operation. Other sustaining capital costs include progressive reclamation costs, starting in Year 2 until the end of the mine life, totalling approximately $35.5M and site closure costs of approximately $24.6M incurring in Year 16. Equipment salvage value, housing sales and the construction camp water treatment plant residual value total approximately ($64.2M).

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21.13

Exclusions

The following items are not included in the Capital Cost Estimate:

•

Inflation and escalation;

•

Costs associated with hedging against currency fluctuations;

•

All taxes, duties and levies;

•

Working capital; and

•

Project financing costs including (interest expense, fees, commissions, etc.).

21.14

Assumptions

•

The use of overburden and NPAG waste rock generated during the mine pre-stripping will be maximized for sourcing backfill material;

•

Other required backfill materials will be available from the nearby borrow pit owned by Rainy River Resources;

•

Bulk earthworks and haulage road construction will be performed by crews assigned to the mine pre-stripping operation;

•

Soil conditions will not require special foundation designs such as piling (as established from geotechnical data and from foundation design recommendations by AMEC);

•

All excavated material will be disposed on-site; and

•

The Project will adhere to the schedule in Section 24.

21.15

Operating Costs Summary

Operating costs were calculated based on open pit and underground mining, processing, refining, transport and general and administrative costs. Salaries were based on a recent salary survey for similar sized projects in northern Ontario. The overall Project operating costs (in CAD$/tonne milled) are shown in Table 21-15. Over the life of the Project, the total operations personnel will peak at approximately 600 employees.

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Table 21-15: Operating Cost Summary


Area	Units	Unit Costs (LOM)
Open Pit Mining (Ore + Waste + Stockpile)^1,2	$/t milled		7.19
Underground Mining³	$/t milled		2.02
Processing	$/t milled		8.65
General & Administration	$/t milled		1.21
Refining and Transportation	$/t milled		0.14
Royalty Payments	$/t milled		0.54

Total	$/t milled		19.75

¹	Open pit mining costs shown here equate to $1.95 per tonne mined.

²	Stockpile reclamation costs of $0.337 per tonne milled are included in the open pit mining costs.

³	Underground mining costs shown here equate to $75.52 per tonne mined, reflecting the weighted average of long hole cut and fill underground mining costs.

Table 21-16 provides a summary of average cash costs per ounce of gold over the first ten years of the Project, as well as over the life of mine. Full cash costs, including silver credits and royalties for the first ten years of operation, total USD $468/oz. Au and USD $544/oz. Au over the life-of-mine. The waste rock costs have been capitalized based on the quantity of material above the average LOM stripping ratio (>3.10). This amounts to a total quantity of 32.6 Mt of waste rock removal which has been capitalized.

Table 21-16: Key Project Operating Costs¹


Area	Initial 10 Year $US/oz. Au			LOM $US/oz. Au
Mining (Open Pit and Underground)²		268			275
Processing		188			258
General and Administrative		28			36
Refining Expenses		4			4
Royalties		18			16
Silver Credit		(38	)		(45	)

Total		468			544

Includes silver credit and royalty payments. Cash costs are calculated by recording the mining cost of stockpiled material in the periods in which the material is processed and revenue recognized, in accordance with the IFRS. During years with high stripping ratios (>3.10), the operating costs associated with rock material have been capitalized.

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The cash costs per ounce of gold vary significantly, depending on the feed grade, mine strip ratio and the amount of stockpiling. The annual fluctuation of operating costs per ounce of gold produced can be seen in Figure 21-2. It should be noted that the overall operating costs per ounce increase quite substantially during the later years due to the processing of stockpile material (with gold grades ranging from 0.3 to 0.6 g/t Au).

LOGO

Figure 21-2: Annual Operating Cash Costs (USD/oz. Au) with Silver Credit

21.15.1

Power and Fuel Costs

The cost of electrical power for the Project was determined to be $0.06/kWh. This includes a reduction for the North Industrial Rebate Program and is a projected 2012 rate based on historical rates from 2006 to 2011.

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21.15.2

Open Pit Mining Operating Costs

Mine operating costs were estimated by BBA using the list of equipment, manpower requirements and mine plan presented in Section 16 of this report. Mining operating costs include the equipment operating cost, the electricity and fuel costs, the salaries, the cost for blasting and other services.

The annual open pit mine unit operating cost is presented in Table 21-17. The average unit operating cost for the mine over the life of the mine is CAD $1.95 per tonne of material mined.

Table 21-17: Open Pit Mine Operating Cost Breakdown


Cost Area	Operating Cost ($/t mined)		%
Hauling		0.61		33
Loading		0.17		9.1
Drilling		0.08		4.3
Support and Auxiliary Equipment		0.30		16
Blasting		0.33		18
Workforce		0.45		24
Services		0.01		0.4

Total Open Pit Operating Cost		1.95		100

Equipment operating costs consist mainly of maintenance costs, which have been estimated by BBA based on Supplier information. Maintenance costs include the costs of repairs, spare parts, consumables, etc., and are compiled on a cost per hour of operation basis for each equipment type.

Diesel fuel is used to operate mine trucks, hydraulic diesel shovels, loaders, dozers and other smaller mine equipment. Fuel consumption was estimated for each year of operation based on equipment specifications, utilization and haulage profile for trucks. A price of $0.85 per litre is used in diesel cost calculations. Electrical power is supplied to the open pit by a power loop and is used to operate the two (2) electric hydraulic shovels. Power consumption was estimated for each year

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of operation based on equipment specifications and equipment utilization. A price of $0.06 per kWh is used for electricity cost calculations.

Blasting costs for ore and waste rock have been estimated based on parameters and powder factors presented in Section 16 of the Feasibility Study report. Blasting unit costs were estimated at $0.29/t for ore and $0.23/t for waste rock, based on an emulsion unit cost of $82.20 per 100 kg. Blasting costs also include accessories and contractor labour costs for mixing, delivering explosives to the blast holes and loading explosives into the blast holes.

Labour requirements have been estimated on an annual basis to support the mine plan developed in this Study. There are a total of 299 employees in the Open Pit; 248 hourly and 51 salaried. Detailed salaried and hourly personnel positions and corresponding headcounts were presented in Section 16 of this report.

Table 21-18 and Table 21-19 present the mine salaried and hourly personnel annual wages, including fringe benefits for the various positions and functions. Salaried and hourly personnel base salaries were provided by Rainy River Resources and are based on local competitive salaries. Benefits were estimated as a percentage of base salary.

Service costs include items such as consultant fees, software licences, an allowance for mine dewatering, dispatch system licence, survey and monitoring equipment and pre-split (the drilling of holes in finished walls prior to blasting to prevent stress cracking).

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Table 21-18: Salaried Mine Personnel Annual Cost (Open Pit)


Mine Salaried Personnel	Annual Cost¹
Mine Superintendent	$	210,000
General Mine Foreman	$	168,750
Mine Shift Foreman	$	156,000
Drill & Blast Foreman	$	156,000
Blaster	$	130,000
Blaster Helper	$	97,500
Dispatcher	$	130,000
Training Foreman	$	130,000
Production / Mine Clerk	$	58,500
Secretary	$	58,500
Maintenance Superintendent	$	210,000
Maintenance General Foreman	$	168,750
Maintenance Planner	$	130,000
Mechanical Engineer	$	140,400
Maintenance Foreman	$	156,000
Maintenance Trainer	$	130,000
Maintenance Clerk	$	58,500
Chief Engineer	$	210,000
Senior Mine Planning Engineer	$	168,750
Open Pit Engineer	$	140,400
Geotechnical Engineer	$	140,400
Blasting Engineer	$	140,400
Mining Engineering Technician	$	97,500
Mine Surveyor	$	91,000
Assistant Surveyor	$	91,000
Chief Geologist	$	210,000
Geologist	$	130,000
Grade Control Geologist	$	130,000
Geology Technician	$	97,500

¹	Annual cost includes bonus and burden.

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Table 21-19: Hourly Mine Personnel Annual Cost (Open Pit)


Mine Hourly Personnel	Annual Cost¹
Shovel Operators	$	84,500
Loader Operators	$	84,500
Haul Truck Operators	$	65,000
Drill Operators	$	71,500
Dozer Operators	$	71,500
Grader Operators	$	71,500
Water Truck Operator/ Snow Plow / Sanding	$	58,500
Other Auxiliary Equipment	$	58,500
General Labour	$	52,000
Janitor	$	45,500
Field Gen Mechanics	$	91,000
Field Welder	$	91,000
Field Electrician	$	91,000
Shovel Mechanics	$	91,000
Shop Electrician	$	91,000
Shop Mechanic	$	91,000
Mechanic Helper	$	58,500
Welder-Machinist	$	91,000
Lube/Service Truck	$	58,500
Electronics Technician	$	58,500
Tool Crib Attendant	$	52,000

¹	Annual cost includes bonus and burden.

Stockpile Reclaim

After depletion of the pit, the mill will be fed at 21,000 tpd with the low-grade stockpiled material. The cost estimate for reclamation is $0.91/tonne of reclaimed ore material ($0.337/t milled LOM) for the transportation of the stockpiled material from the stockpile to the crusher. This cost includes the equipment maintenance, fuel, and the labour required for maintenance and operations.

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21.15.3

Underground Mining Operating Costs

The operating costs for the RRU were estimated from first principles, using quotes for equipment operating costs and for major consumables such as explosives, ground support, and cement. Labour costs were estimated in consultation with Rainy River and BBA. Also included in the operating costs are the following:

•

Services such as setting up a drawpoint for remote mucking, definition diamond drilling, and installing air and water pipes;

•

Cemented backfill of the primary and longitudinal stopes, and rockfill for the secondary and CAF stopes;

•

Operating development such as the excavation of the drawpoints for the longhole stopes and attack ramps for the CAF stopes; and

•

Fixed costs such as power for the underground (e.g., ventilation fans and pumps), and propane for the heaters.

Each item in the table contains the cost of consumables and equipment required to complete the activity. For example, the drilling activity includes the cost of rods, bits and adapters required to drill the hole, the cost to maintain the longhole drill, and the cost of the fuel to move the longhole drill from one stope to the next.

The ore haulage costs include the truck operating cost per hour and fuel consumption to haul ore out of the mine. The cost of the return trip is allocated either to ore haulage or backfill, depending on the destination of the returning truck. It is estimated that the majority of the development waste will be used to backfill the CAF areas.

General mine expenses have not been included in this estimate. They have been estimated by BBA for the open pit and underground.

The life-of-mine operating cost of the RRU is estimated to be $75.52/tonne and a cost breakdown is presented in Table 21-20.

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Table 21-20: Breakdown of the Rainy River Underground Operating Cost by Activity


Activity	Longhole Operating Cost ($/tonne)		CAF Operating Cost ($/tonne)
Drilling		1.52		1.32
Blasting		1.37		5.26
Mucking		2.78		1.31
Ground Support		0.25		9.30
Services Provision		0.69		3.32
Ore Haulage		1.58		2.04
Backfill		10.33		5.11
Operating Development		5.07		10.50

Subtotal		23.59		38.14

Labour		37.73
Fixed		11.91

Total		75.52

Table 21-21 shows the operating cost for the life of the RRU. It varies from a high of $115/tonne during initial production to a low of $65/tonne during steady state production. It is estimated that the underground mining at the RRU will cost $234M over the life of the mine.

Table 21-21: Rainy River Underground Yearly and Total Operating Cost (‘000)


Year	CAF Consum.		CAF OPEX Develop.		Longhole Consum.		Longhole OPEX Develop.		Fixed		Labour		Total OPEX		Total OPEX ($/t)
2016 (Q3)		—		—		—		—		—		—		—		—
2016 (Q4)		—		—		—		—		—		—		—		—
2017 (Q1)		—		—		—		—		—		—		—		—
2017 (Q2)		—		—		—		—		—		—		—		—
2017 (Q3)		—		—		—		—		—		—		—		—
2017 (Q4)		—		—		—		—		—		—		—		—
2018 (H1)		—		—		—		—		—		—		—		—
2018 (H2)		—		—		—		—		—		—		—		—
2019		—		925		2,630		1,013		3,000		9,581		17,149		115.78
2020		1,325		616		3,718		74		3,181		9,662		18,576		74.56
2021		2,629		700		4,960		5,448		3,560		17,295		34,592		84.09
2022		2,629		945		4,178		2,889		3,819		14,398		28,858		85.19

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Year	CAF Consum.			CAF OPEX Develop.			Longhole Consum.			Longhole OPEX Develop.			Fixed			Labour			Total OPEX		Total OPEX ($/t)
2023		2,616			939			4,640			2,230			3,819			13,648			27,892		77.61
2024		2,782			1,261			4,750			282			3,923			10,881			23,879		66.20
2025		2,788			507			4,893			—			3,923			11,583			23,694		64.70
2026		2,491			668			4,933			—			3,923			11,583			23,598		65.99
2027		6,320			242			4,981			—			3,923			10,703			20,481		70.10
2028		—			—			4,059			51			3,923			7,882			15,915		72.57

Total		17,892			6,803			43,742			11,987			36,994			11,7216			234,634		75.52

Percentage		8	%		3	%		18	%		5	%		16	%		50	%

21.15.4

Process Plant Operating Costs

Process plant operating costs were calculated for 16 years of operation. The operating costs are based on metallurgical testwork, the mine plan, a recent salary survey, literature, and recent supplier quotations. The average life-of-mine processing operating costs were determined to be $8.65/tonne milled at approximately 21,000 tpd. The yearly tonnages from the mine plan vary and were used to obtain the operating costs. As expected, operating costs for the first year are higher as the plant is still ramping up to full production. The average operating cost includes reagents, consumables, grinding media, personnel (including contractors), electrical power, propane and maintenance parts. The consumables include spare parts, grinding media and liner and screen components. The fluctuations of processing costs over the mine life are presented in Table 21-22.

Table 21-22: Process Operating Costs


Year	$/tonne milled
1		10.00
1-10		8.67
2-16		8.62
LOM		8.65

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Reagents and Consumables

The reagent consumptions were estimated based on testwork, industrial references, literature and assumed operational practice.

Suppliers provided quotes for various yearly consumed spare parts. The annual cost for grinding media for the SAG and ball mill were estimated based on the expected media consumption (g/kWh) and the cost per tonne of steel media. The respective power required for the SAG and ball mill were calculated based on the material hardness (Axb value) for the SAG mill and the BWi of the ball mill. Liner and screen components were calculated based on supplier quotes for the estimated parts cost and their expected wear life, provided in months.

The individual reagent costs ($/t reagent) and media costs ($/tonne steel media) were established through supplier communication and comparison with prices at reference sites.

Process Plant Personnel

The personnel costs incorporate requirements for plant management, metallurgy, operations, maintenance, site services, assay lab and contractor allowance. The individual personnel are divided into their respective positions and their salaries were provided by the client or taken from the 2012 Canadian Mines Salary Survey from similar sized gold operations located in Ontario. It was assumed that contractors would be used for crusher and liner changes.

There are a total of 89 employees accounted for in the process operating costs, 72 employees in the process plant and 17 employees in the assay lab.

Electricity and Propane

The propane consumption includes requirements for the crusher, concentrator, process and boiler. The cost of propane ($0.50/L) was provided by a reference site.

The process plant energy consumption was estimated from the equipment running loads and the SAG and ball mill grinding energy requirements. Various factors (efficiency, load, diversity and annual factors) were applied to adjust for equipment motor efficiency, the power used versus

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installed, the synchronous operation of equipment and average on-line time in the plant and throughout the year.

The SAG mill specific energy (kWh/t) was estimated from the analyzed relationship derived from testwork between the Axb value and the SAG motor input specific energy as determined by JK SimMet. The ball mill specific energy (kWh/t) was calculated from the BWI and the Bond formula, assuming the ball mill will grind the rock from 2,400 µm to 75 µm. The Axb and BWI values were estimated for each year based on testwork and the mine plan. The specific energies were converted to an annual power demand (GWh) based on efficiency factors and the varying milled tonnage per year.

A breakdown of the average processing operating costs can be seen in Table 21-23.

Table 21-23: Yearly Average Processing Operating Cost Breakdown


Cost Area	Operating Cost ($/t milled)		%
Reagents		2.05		23.7
Spare Parts/Maintenance		0.56		6.4
Liners and Screen Components		0.46		5.4
Grinding Media		1.79		20.7
Personnel		1.06		12.3
Electrical Power		2.51		29.0
Propane		0.21		2.4

Total		8.65		100

It can be seen that the main cost areas for the process plant are the electrical power, grinding media and reagents. The majority of the reagent costs are associated with cyanide leaching and cyanide destruction. The personnel costs represent approximately 12% of the processing costs.

21.15.5

General and Administrative Costs

General and Administrative (G&A) costs are expenses not directly related to the production of goods and encompass items not included in mining, processing, refining and transportation costs. These costs are based on the client’s recommendations, similar sized operations, and BBA’s in-house database.

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The G&A costs were calculated for 16 years of operation and with an average cost of $1.21/tonne milled. This cost includes:

•

Human Resources;

•

Site Administration, Insurance and Management;

•

Infrastructure Power;

•

Health and Safety;

•

Assay Laboratory supplies;

•

Environmental costs;

•

G&A Personnel;

•

Information Technology (IT);

•

Insurance; and

•

Training.

The labour costs account for the 26 employees required for G&A, including 16 staff and 10 hourly employees. In general, the management and administrative staff will work 40 hours per week on a day shift. Security and warehousing personnel will work a 12-hour shift per day to support the 24-hours per day operations requirements. The labour cost represents 23% of the G&A cost. Infrastructure electricity accounts for approximately 26% and site administration, maintenance and insurance represent approximately 32% of the G&A costs. The breakdown is shown in Table 21-24.

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Table 21-24: Average General and Administrative Costs


Cost Area	Operating Cost ($/t milled)		%
Infrastructure Electricity		0.31		26.0
Site Admin., Insurance and Maintenance		0.38		31.7
Assay Laboratory Consumables		0.02		1.4
Health and Safety		0.07		6.2
Environment		0.03		2.7
Human Resources		0.06		4.8
IT and Telecommunications		0.06		4.7
Personnel		0.27		22.5

Total		1.21		100

21.15.6

Royalties

The annual royalty costs are based on the conceptual mine design and production profile, along with the terms of the individual royalty agreements, which in turn relate to a limited portion of the reserves. Royalty costs are based on a gold price of USD $1,400 per ounce gold and USD $25 per ounce silver. Over the life of the Project, approximately $63.2M in royalties is expected to be paid. These royalties include an annual 10% net profit interest payable to the royalty holder once all pre-production capital expenditures have been recovered, and the prorated share of operating and sustaining costs for the property have been deducted.

21.15.7

Transportation and Refining

A weekly shipment of doré bars containing approximately 38% gold and 62% silver will be transported to a refinery. A flat rate transportation cost will be incurred by the refinery in addition to a cost by weight and a variable liability fee. A treatment cost per troy ounce of material shipped to the refinery will also be charged. Rainy River will be paid for a set recovery of the assayed content. Over the life of the mine, a transport and refining cost of 0.14 $/t milled is estimated and this is based on a budgetary quotation obtained from a North American gold refinery.

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22.

ECONOMIC ANALYSIS

A financial analysis for the Rainy River Gold Project was carried out using a discounted cash flow approach on a pre-tax and after-tax basis. The internal rate of return (IRR) on total investment was calculated based on 100% equity financing. The Net Present Value (NPV) was calculated from the cash flow generated by the project based on a discount rate of 5%. The payback period based on the undiscounted annual cash flow of the project was also indicated as a financial measure. Furthermore a sensitivity analysis was also performed for the after-tax base case to assess the impact of variations of the project capital costs, annual operating costs, price of gold/silver and recovery of gold/silver.

The economic analysis presented in this section contains forward-looking information with regards to the mineral reserve estimates, commodity prices, exchange rates, proposed mine production plan, projected recovery rates and processing costs, infrastructure construction costs and schedule. The results of the economic analysis and are 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.

22.1

Assumptions and Basis

The Economic Analysis was performed using the following assumptions and basis:

•

The Project Executive Schedule developed in the FS (Chapter 24), considering key project milestones;

•

The Financial Analysis was performed for the entire mineral reserve estimated in this Study;

•

Commercial production start-up is scheduled to begin in third quarter (Q3) of 2016. The first full year of production is therefore 2017. Operations are estimated to span a period of approximately 16 years;

•

The base case gold and silver prices are USD $1,400/oz. and USD $25/oz., respectively;

•

The United States to Canadian dollar exchange rate has been assumed to be USD $1.00:CAD $1.00 during the first two (2) preproduction years and USD $1.00:CAD $1.07 during operations;

•

All cost and sales estimates are in constant Q4 2012 Canadian dollars with no inflation or escalation factors taken into account;

•

All gold and silver is sold in the same year it is produced;

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•

The financial analysis does not include working capital for the period between commissioning and first metal sales, or closure plan bonding requirements. It is assumed that these items are covered by project financing or by a negotiated loan facility; and

•

Final rehabilitation and closure costs will be incurred after production Year 16.

This Financial Analysis was performed by BBA on a pre-tax basis. Rainy River Resources Management provided the after-tax economic evaluation of the project, which was prepared with the assistance of an external tax consultant. The general assumptions used for this financial model are summarized in Table 22-1.

Table 22-1: Financial Model Criteria


Description	Value		Unit
Daily Milling Rate		21,000	tpd
Open Pit Mining¹		7.19	$/t milled (LOM)
Underground Mining		2.02	$/t milled (LOM)
Processing		8.65	$/t milled (LOM)
General and Administration		1.21	$/t milled (LOM)
Refining Expenses		0.14	$/t milled (LOM)
Royalties		0.54	$/t milled (LOM)
Gold Recovery		90.4	% (LOM)
Silver Recovery		64.1	% (LOM)
Long Term Gold Price		1400	USD$/oz.
Long Term Silver Price		25	USD$/oz.
Pre-Production Exchange Rate		1.00	CAD$/USD$
Long Term Exchange Rate		1.07	CAD$/USD$
Discount Rate		5	%
Open Pit - Initial Capital Cost²		713.3	$M
Open Pit - Sustaining Capital Cost³		321.9	$M
Underground – Development Phase Capital Cost³		67.8	$M
Underground – Sustaining Capital Cost³		94.6	$M

¹	Stockpile reclamation costs of $0.337 per tonne milled are included in open pit mining costs.

²	Includes pre-production process plant and infrastructure capital costs.

³	Funded by internal cash flows. Includes process plant and infrastructure sustaining capital costs.

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22.2

Royalties

The annual royalty costs were calculated by RRR and are based on the conceptual open pit and underground mine design and production profiles developed by BBA and Golder, along with the terms of the individual royalty agreements, which in turn relate to a limited portion of the reserves. Royalty costs are based on a gold price of USD$1,400 per ounce of gold and USD$25 per ounce of silver. Over the life of the Project approximately $63.2M in royalties is expected to be paid based on the base case metal prices and project assumptions. These royalties include an annual 10% net profit interest payable to the royalty holder once all pre-production capital expenditures have been recovered, and the prorated share of operating and sustaining costs for the property have been deducted.

22.3

Taxation

The RRGP is subject to three (3) levels of taxation, including federal income tax, provincial income tax and provincial mining taxes. Rainy River compiled the taxation calculations for the RRGP with assistance from third-party taxation experts, and the calculation includes the impact of changes announced in the March 2013 federal budget. Following are highlights of the principal elements of taxation that have been reflected in the tax calculations for the RRGP. The information below was provided by Rainy River, and was not verified by BBA:

•

Income tax is payable to the federal government of Canada, pursuant to the Income Tax Act (Canada). The applicable federal income tax rate is 15% of taxable income;

•

Income tax is payable to the province of Ontario at a tax rate of 10% of taxable income, which includes the manufacturing and processing tax credit. Ontario income tax is administered by the Canada Revenue Agency and, since 2008, Ontario’s definition of taxable income is fully harmonized with the federal definition;

•

Ontario Mining Tax (“OMT”) is levied at a rate of 10% on taxable profit in excess of $500,000 derived from mining operations in Ontario. OMT is deductible in calculating federal income tax and a similar resource allowance is available as a deduction in calculating Ontario income tax. OMT is not affected by harmonization; accordingly, it is administered provincially by Ontario;

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•

Corporate Minimum Tax (“CMT”) is applied to Ontario corporations that have at least $100M in gross revenues and at least $50M in total assets, at a rate of 2.7%. The CMT is payable only to the extent that it exceeds the regular Ontario corporate income tax liability, and any CMT that is paid is creditable against regular Ontario corporate income tax payable in the next 20 years; and

•

The combined effect on the Project of the three (3) levels of taxation, including the elements described above, is a cumulative effective tax rate of 28.5%, based on Project totals.

The tax calculations are underpinned by the following key assumptions:

•

The Project is held 100% by a corporate entity and the after-tax analysis does not attempt to reflect any future changes in corporate structure or property ownership;

•

Assumes 100% equity financing and therefore does not consider interest and financing expenses;

•

Payments projected relating to NSR or NPI royalties, as applicable, are allowed as a deduction for federal and provincial income tax purposes, but are added back for provincial mining tax purposes; and

•

Actual taxes payable will be affected by corporate activities, and current and future tax benefits have not been considered.

22.4

Financial Analysis Summary

A discount rate of 5% was applied to the cash flow to derive the project’s NPV on a pre-tax and after-tax basis. The summary of the financial evaluation for the Project’s base case is presented in Table 22-2.

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Table 22-2: Financial Analysis Summary (Pre-tax and After-tax)


Description		Base Case			Units
	Net Present Value (5% disc)		1,296			$M
	Internal Rate of Return		27.8	%		%
	Simple Payback Period		3.1			Years
	Net Present Value (5% disc)		931			$M
	Internal Rate of Return		23.7			%
	Simple Payback Period		3.2			Years

The pre-tax base case financial model resulted in an internal rate of return of 27.8% and a NPV of $1,296M with a discount rate of 5%. The simple payback period is 3.1 years. On an after-tax basis, the base case financial model resulted in an internal rate of return of 23.7% and a NPV of $931M with a discount rate of 5%. The simple payback period is 3.2 years.

The summary of the Rainy River discounted cash flow financial model (pre-tax and after tax) is presented in Table 22-3.

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Table 22-3: Rainy River Financial Model Summary

LOGO

Year -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 Total

Production Summary

Total Tonnes Mined OP & UG (Mt) 0.0 0.0 6.0 14.1 13.0 11.8 13.0 13.3 12.5 11.7 11.3 7.7 0.8 0.3 0.2 0.0 0.0 0.0 116

Total Tonnes Milled (Mt) 0.0 0.0 2.9 7.7 7.7 7.7 7.7 7.7 7.7 7.7 7.7 7.7 7.7 7.7 7.7 7.7 7.7 6.2 116

Blended Mill Head Grade Au (g/t) 0.0 0.0 1.2 1.3 1.4 1.8 1.7 1.5 1.3 1.3 1.4 1.6 0.6 0.5 0.5 0.3 0.3 0.5 1.1

Blended Mill Head Grade Ag (g/t) 0.0 0.0 3.1 3.0 3.9 2.1 2.4 2.7 3.3 5.0 3.8 2.4 2.3 2.3 1.5 1.9 2.1 2.3 2.8

Gold Recovery (%) 0.0% 0.0% 90.9% 91.5% 91.6% 92.9% 92.6% 92.2% 91.5% 91.4% 91.6% 92.4% 86.8% 85.2% 84.4% 79.8% 78.8% 74.4% 90.4%

Silver Recovery (%) 0.0% 0.0% 64.0% 64.2% 63.0% 65.5% 65.0% 64.6% 63.7% 61.3% 63.0% 65.1% 65.2% 65.3% 66.4% 65.7% 65.5% 64.9% 64.1%

Gold (‘000s oz.) 0.0 0.0 99.8 295.7 306.2 403.8 383.3 349.7 298.8 294.3 305.4 359.6 136.4 112.0 102.0 63.1 57.8 77.0 3,645

Silver (‘000s oz.) 0.0 0.0 185.3 472.5 598.1 340.0 388.8 434.1 525.6 760.8 597.5 386.9 371.5 364.9 244.6 311.6 339.2 293.7 6,615

Revenue

Exchange Rate ($CAD:$USD) 1.00 1.00 1.07 1.07 1.07 1.07 1.07 1.07 1.07 1.07 1.07 1.07 1.07 1.07 1.07 1.07 1.07 1.07

Gross Revenue ($M) 0.0 0.0 154.5 455.6 474.7 614.0 584.6 535.4 461.7 461.2 473.4 549.1 214.3 177.6 159.4 102.8 95.7 123.2 5,637

Operating Expenditures

Mining ($M) 0.0 0.0 34.6 81.8 80.5 92.7 112.9 133.6 125.9 119.0 119.2 68.3 32.8 27.2 22.7 7.0 7.0 5.6 1,071

Processing ($M) 0.0 0.0 29.2 67.6 67.1 67.6 67.4 67.1 67.4 67.8 67.0 67.5 66.8 66.5 66.7 66.8 66.2 53.9 1,023

General and Administration ($M) 0.0 0.0 4.9 9.6 9.6 9.6 9.6 9.6 9.6 9.6 9.6 9.6 9.6 9.1 9.1 7.5 7.5 6.0 141

Royalty Payments ($M) 0.0 0.0 0.00 0.00 0.00 0.00 9.56 13.66 5.17 4.61 5.78 21.60 2.44 0.31 0.15 0.00 0.00 0.00 63

Capital Expenditures

Pre-Production ($M) 139.0 272.4 134.2 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 546

Sustaining ($M)1 0.0 0.0 38.0 77.2 69.1 82.1 63.3 43.4 52.3 50.4 23.3 7.7 -25.7 4.7 1.5 0.9 0.9 -4.9 484

Indirect Costs

Indirects ($M) 28.2 56.4 28.2 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 113

Contingency ($M) 13.8 27.5 13.8 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 55

Pre-Tax Cash Flow

Pre-Tax Cash flow ($M) Cumulative Pre-Tax Cash flow ($M) After-Tax Cash Flow After-Tax Cash flow ($M) Cumulative After-Tax. Cash flow ($M) Pre-Tax NPV @ 5% ($M) Pre-Tax IRR After-Tax NPV @ 5% ($M) After-Tax IRR (180.9) (180.9) (180.9) (180.9) 1,295.9 27.8% 931.2 23.7% (356.3) (537.2) (356.3) (537.2) (128.5) (665.7) (128.5) (665.7) 219.3 (446.4) 211.5 (454.2) 248.3 (198.1) 238.8 (215.4) 361.9 163.8 333.3 117.9 321.7 485.5 238.0 355.9 268.0 753.5 201.0 556.8 201.3 954.8 145.2 702.0 209.8 1,164.6 146.2 848.2 248.5 1,413.1 173.4 1,021.6 374.2 1,787.3 265.7 1,287.3 128.3 1,915.6 106.3 1,393.7 69.7 1,985.3 56.3 1,450.0 59.3 2,044.6 48.5 1,498.5 20.7 2,065.3 20.1 1,518.6 14.1 2,079.4 14.1 1,532.7 62.6 2,142.0 53.3 1,586.0 2,142 2,142 1,586 1,586

1 Includes underground development phase and sustaining capital, in addition to open pit sustaining capital.

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Figure 22-1 shows the projects cumulative cash flows projected for the life of the mine on a pre-tax and after-tax basis.

LOGO

Figure 22-1: Life-of-Mine Cash Flow Projection (Pre-tax and After-tax, discount rate: 5%)

22.5

Financial Model Sensitivity Analysis

A financial sensitivity analysis was conducted on the base case cash flow net present value and internal rate of return of the Project, based on the following variables: capital expenditures, operational expenditures, metal recoveries and metal prices. The pre-tax and after tax results, based on various metal prices and exchange rates, are summarized below in Table 22-4. Please note that the royalty payments were assumed to be constant (equivalent to the base case) for the sensitivity analysis.

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Table 22-4: Sensitivity Results for Metal Price and Exchange Rate Variations


Gold Price (USD$ / ounce)		Silver Price (USD$ / ounce)		Exchange Rate (USD$:CAD$)		NPV at 5% Discount Rate (CDN$ million)				IRR (%)				Payback Period (years)
Gold Price (USD$ / ounce)		Silver Price (USD$ / ounce)		Exchange Rate (USD$:CAD$)		Pre-Tax		After- Tax		Pre-Tax		After- Tax		Pre-Tax		After- Tax
$	1,250	$	25		0.91	$	1,002	$	721		23.5		19.9		3.4		3.5
$	1,400	$	25		0.93	$	1,296	$	931		27.8		23.7		3.1		3.2
$	1,600	$	30		0.97	$	1,674	$	1,191		32.8		27.9		2.7		2.8
$	1,800	$	35		1.00	$	2,059	$	1,469		37.6		32.1		2.4		2.5

A summary of the after-tax sensitivity analysis for project capital costs, operating costs and metal recovery is summarized in the following table.

Table 22-5: Sensitivity Results for CAPEX, OPEX and Metal Recovery Variations (After-tax)


CAPEX	LOM CAPEX ($)						NPV @ 5% ($)		IRR (%)
120%		1,436,813,959						771,760,913		18.3	%
110%		1,315,378,399						851,683,336		20.8	%
100% (Base Case)		1,197,657,116						931,204,128		23.7	%
90%		1,083,391,245						1,005,877,517		26.9	%
80%		972,321,922						1,076,511,849		30.6	%

OPEX	LOM OPEX ($)						NPV @ 5% ($)		IRR (%)
120%		2,770,118,754						778,864,926		20.8	%
110%		2,532,361,596						855,795,920		22.2	%
100% (Base Case)		2,297,414,854						931,204,128		23.7	%
90%		2,065,278,528						1,004,749,832		25.0	%
80%		1,835,952,618						1,074,791,933		26.3	%

Recovery	Au Recovery (%)			Ag Recovery (%)			NPV @ 5% ($)		IRR (%)
105%		95.9	%		68.0	%		1,013,749,897		25.0	%
103%		93.1	%		66.1	%		972,976,205		24.4	%
100% (Base Case)		90.4	%		64.1	%		931,204,128		23.7	%
95%		81.6	%		57.9	%		791,566,092		21.2	%
90%		73.2	%		51.9	%		651,771,015		18.7	%

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The graphical representations of the financial sensitivity analysis summary are depicted below in Figure 22-2 and Figure 22-3. It can be seen that the Project financials are most sensitive to metal price and overall recoveries. Overall, the NPV of the RRGP is positive and the project is robust under all of the sensitivity conditions when analyzed individually.

LOGO

Figure 22-2: Sensitivity of the Net Present Value (After-tax) to Financial Variables

LOGO

Figure 22-3: Sensitivity of the Internal Rate of Return (After-tax) to Financial Variables

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23.

ADJACENT PROPERTIES

Introduction

Seven (7) properties in the exploration stage are located adjacent to or near the Rainy River Gold Project property. Bayfield Ventures Corp. holds three (3) of the properties, known as the B Block, C Block and Burns Block. Coventry Resources Inc. holds three (3) of the properties, known as the Pattullo, Nelles and Blue properties. King’s Bay Gold Corp. holds the seventh property, being a single continuous land package contiguous with the most northerly portion of the Rainy River Gold Project property. The closest Canadian operating mine is the Lac des Iles; palladium, nickel, gold and copper mine, owned by North American Palladium Ltd., located 90 km northwest of Thunder Bay and nearly 400 km northeast of the Rainy River Gold Project.

The following information was obtained from public sources, and neither Rainy River nor BBA has verified its accuracy. The information provided in the following section is not a reflection of the mineralization of the Rainy River Gold Project property.

Bayfield Ventures Corp.

Bayfield Ventures Corp. (“Bayfield”), exploring for gold and silver, owns 100% interest in mineral rights to three (3) properties adjacent to the Rainy River Gold Project property. These consist of the B Block, C Block and the Burns Block as shown in Figure 23-1.

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LOGO

Figure 23-1: Bayfield Ventures Corp. Properties

(Source: www.bayfieldventures.com/i/pdf.BYVRRArea.pdf)

The Burns Block, which is Bayfield’s gold-silver flagship project, lies directly east of the ODM Zone and west of the newly discovered Intrepid Zone, and is completely surrounded by properties either owned by, or optioned to Rainy River Resources (both surface and mineral rights). In 2012, Rainy River also purchased a 100% interest in the surface rights to the Burns Block. Although several significant gold mineralized intersections on the Burns Block have been reported by Bayfield, to Rainy River’s knowledge the Burns Block does not contain any “mineral resource” (as that term is defined by the Canadian Institute of Mining, Metallurgy and Petroleum, and incorporated by reference in National Instrument 43-101), nor has there been a 43-101 compliant technical report completed for the Burns Block. Bayfield, however, did publish a NI 43-101 compliant technical report in March 2007 on a property 3 km southwest of Rainy River’s proposed open pit over which Bayfield held an option. That option was subsequently allowed to lapse and the property has been returned to the original mineral rights owner. No significant drill results have been reported by Bayfield for the B Block or the C Block.

Coventry Resources Inc.

Coventry Resources Inc. (“Coventry”) has a gold exploration project in the Rainy River District comprised of three (3) properties to the west of the western perimeter of the Rainy River Gold Project, known as the Pattullo, Nelles and Blue properties. These are shown in Figure 23-2.

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LOGO

Figure 23-2: Coventry Resources Properties

(Source: http://www.coventryres.com/s/RainyRiver.asp)

Coventry holds patented claims optioned by landholders, staked unpatented claims and unpatented claims optioned by third parties. Within the next seven (7) years, Coventry Resources is entitled to earn a 100% interest in the mineral rights on all leased areas (132.7 km2).

In a news release issued on January 28, 2013, Coventry indicated that it was starting a 6,000 m drilling program on the aforementioned properties to further evaluate the gold and base metal anomalies discovered in its previous drilling program. A news release issued on February 18, 2013 indicated that Coventry had completed geological mapping, geochemical water sampling and chemical analyses on 181 drill holes. At present, Coventry has not started any studies or preliminary economic assessments on their properties.

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King’s Bay Gold Corp.

King’s Bay Gold Corp. (“King’s Bay”), exploring for gold, owns a 100% interest (both surface and mineral rights) and 18 Menary claims (1,728 hectares), shown in Figure 23-3. The Menary property is located approximately 15 km northeast of the Rainy River’s proposed open pit. The Menary property is contiguous with Rainy River’s Off Lake properties.

LOGO

Figure 23-3: King’s Bay Gold Menary Property

(Source: www.kingsbaygold.com/wp-content/uploads/2011/05/menary-twp-claim-map-2.pdf)

In 2010 and 2011, 33 BQ diamond drill holes were drilled by King’s Bay. In 2011, a shear zone was discovered and channel samples were taken. King’s Bay is considering conducting further stripping and drilling, mapping and sampling. No National Instrument 43-101 compliant technical reports have been issued for the Menary property to date.

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24.

OTHER RELEVANT DATA AND INFORMATION

24.1

Project Execution

The Project Execution Strategy was developed for the Rainy River Feasibility Study (FS) based on the latest Feasibility Study information available and best practices. It is intended to describe the strategy for moving forward on engineering, procurement, construction and environmental activities. The Execution Strategy ensures that the core elements of Rainy River Resources Corporate Objectives regarding the inter-relationship between the stakeholders, community, First Nations and project development are maintained from the mineral exploration stage through the construction phase.

The core elements of the Rainy River Resources Corporate Objectives integrate the following commitments into the strategy:

•

Human Rights;

•

Project Due Diligence and Pre-Engagement;

•

Community and Aboriginal Engagement and Enhancement;

•

Human Resource Development;

•

Environmental Integrity and Performance, and

•

Health and Safety Performance.

24.2

Health, Safety, Environmental and Security

A fully integrated Health, Safety and Environmental (“HSE”) program will be implemented to help achieve a “zero-harm” goal. HSE practices include: alignment with site contractors on safety training, occupational health and hygiene, hazard and risk awareness, safe systems of work, and job safety analysis. A 24-hour staffed site security program will be supplied during the initial field mobilization in 2014.

24.3

Hazardous Waste Management

Specific procedures for waste management and spill response will be implemented for the construction period. These procedures will cover compliance, auditing and reporting requirements.

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Procedures regarding on-going cleanup and rubbish removal, as well as safe handling, storage and disposal of batteries, fuels, oil and hazardous materials, will be established and observed for the duration of the construction phase. Waste will be recycled to the extent feasible. Ongoing dust suppression and rain water management programs will also be established and observed for the duration of the construction phase. Specific procedures and storage areas will be designated for construction waste prior to recycling or removal from the plant. Solid waste will be disposed of in designated pits.

24.4

Execution Strategy

Under the direction of a Construction Management (“CM”) team, field construction contractors will commence work once engineering tasks are well advanced and long lead times for the delivery of major equipment are confirmed.

Basic engineering and detailed engineering begins in the first half of 2013. Construction work packages will be issued on a fixed-price basis dependent on level of engineering progress. Otherwise, construction work packages will be based on a unit price structure incorporating calculated quantities and an estimated budget or target price. Contract packages will be designed for cost and schedule efficiency.

24.5

Management Procedures

The Project Team (the “Team”) including the Owner, engineers and construction managers are responsible for bringing the Project in on time and within budget. Figure 24-1 shows the Project Management Organization Chart envisaged as the Project moves into detailed engineering, procurement and construction.

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LOGO

Figure 24-1: Project Management Organization Chart

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24.6

Project Scheduling

During the Feasibility Study phase, a detailed project schedule was produced which will become the ‘Baseline Schedule’. The overall Project Schedule (“Schedule”) identifies the preferred critical sequences and target milestone dates that need to be managed for the Project to be executed successfully. The detailed schedules track the planned and actual progress throughout the duration of the Project using information provided by the engineering groups, contractors, suppliers, the field management staff and the Owner.

The 23-month project construction duration assumes commencement of field activities in August 2014 to mechanical completion in June 2016.

The August start schedule specifically takes into account 2014/2015 winter works for activities such as bulk earthworks and concrete placement which could have significant cost impact. These activities were deferred to 2015 without impact to the Mechanical Completion date to mitigate such costs.

The Feasibility Study Project Schedule reflects the Environmental Assessment approval and permits in place to enable commencement of construction activities in August 2014. Detailed engineering would commence in 2013 with substantial completion in third quarter (Q3) of 2014, which will facilitate the Project Team’s ability to maximize Lump Sum Construction Contracts.

The Project milestones are summarized in Figure 24-2.

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LOGO

Figure 24-2: Project Milestones

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After acceptance of the Feasibility Study and the associated trade-off studies, the implementation of the Project Schedule will begin by ordering capital equipment items critical to engineering design. At the same time, detail engineering and remaining basic engineering for the plant and infrastructure will start.

24.7

Procurement and Contracts

Purchasing and Expediting Strategy

The EPCM procurement group will provide capital equipment procurement, supplier drawing expediting and coordinate equipment inspection. The procurement group will manage the bidding cycle for equipment and materials to be supplied by the Owner to the contractors. Standard procurement terms and conditions approved for the Project will be utilized for all equipment and materials Purchase Orders. Suppliers will be selected based on location, quality, price, delivery and support service.

24.8

Site Development

Highway 600

Construction of Highway 600 and the East Access Road will be included in the major civil contracts.

230 kV Overhead Line

The construction of the 230 kV power line will be subcontracted to a specialized contractor on a design-build basis.

24.9

Construction

24.9.1

Construction Management Responsibilities

The CM group will be responsible for the management of the construction site. The Construction Manager will be responsible for effectively planning, organizing, and managing the construction quality, safety, budget, and scheduling objectives of the Project.

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The CM Field Engineering Team will employ independent Quality Assurance specialists, qualified to CSA, to ensure the implementation and success of the contractor’s quality control.

24.9.2

Construction Power

The contractors present in the initial phase of the construction will be expected to be self-sufficient in terms of construction electrical power. The main substation at the site and adjoining 230 kV OHL will be constructed by the end of 2014. It is anticipated that 230 kV power will be available in January 2015. Pre-commissioning and commissioning of electrical distribution power lines and equipment will commence thereafter. From February 2015, this will supply power to all mine equipment and peak construction power loads for the balance of the construction phase.

The emergency power generators planned for the primary crusher and the process plant will be purchased early and used for operations of the remote water pump house at the Pinewood/McCallum junction. This pump house has to be operational early in the construction schedule in order to pump fresh water to fill up the WMP. The water stored in the WMP is required for plant start-up.

24.9.3

Construction Labour Requirement

The Schedule has been based on a 70-hour work week with some double-shifts, as required. Crew rotations are planned to be three (3) weeks on site and one (1) week off site.

Approximately 1,330,000 man-hours of direct construction labour are anticipated during project construction, excluding mine pre-development and engineering. Construction manpower is expected to peak at approximately 400 direct construction workers on site, as shown in Figure 24-3. The double-peak curve is a reflection of the slowdown of construction work planned for the winter of 2015. The slowdown reflects the project schedule objective, while reducing the expenses related to winter concrete work.

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LOGO

Figure 24-3: Monthly Construction Manpower Graph

24.10

Process Facilities

24.10.1

Critical Path and Installation Methodology

The Schedule has been presented in association with the established Project Work Breakdown Structure (“WBS”) which defines the elements of project scope, each of which can stand alone with estimate, cost, schedule and accountability.

There is one critical path where there is zero float. It currently runs through the mechanical completion of the SAG mill.

Primary Crusher Installation

Once the foundations for the primary crusher are completed, the MSE retaining walls will be constructed concurrently with the structural and general backfill, as well as the mine truck ramp. The mine truck ramp will be constructed using mine overburden and waste rock which will be

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delivered to the crusher area via the mine trucks. This work will be completed by the fourth quarter (Q4) of 2015.

Mill Installations Including Operating Floors

It has been recommended to install the complete operating floor before the mills are installed. This step will eliminate the need to construct temporary platforms around the mill foundations to enable surveying and assembly of the various shells and head sections.

24.11

Tailings Management Area Earthworks

The TMA will be constructed in staged lifts throughout the mine life. Construction of the TMA and Water Management Pond (WMP) will commence two (2) years prior to mill start-up to ensure sufficient water collection from spring freshets and water sources for use in the mill process.

Construction is constrained by work that restricts the placement of certain general and rock fills to the warm and dry months of the year. The phased completion of the TMA has been scheduled accordingly, governed in part by the engineer’s quality specifications which will determine the “no-build” restrictions during the winter.

Initial construction of the TMA will begin with the water management pond and the southwest starter dam. Materials for construction will be supplied from the open pit pre-stripping development. An on-site quarry will also be used to supply appropriately sized aggregates.

24.12

Commissioning

The EPCM team will be responsible for the installation of facilities, with the exception of all mining activities until mechanical completion.

The Sequence of System Commissioning is vital to shifting the construction schedule from general area completion to more specific system completion to suit the commissioning and start-up of the entire facility.

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During the latter part of engineering, the Owner will develop a commissioning plan in conjunction with the construction management forces. The systems will be identified and scheduled for delivery by priority. Packages will be assembled for each system that must be commissioned to include all sign-off and test documentation, drawings and supplier information.

As the various systems are completed and determined by the Construction Management team to be free of deficiencies that would prevent safe operation, they will be transferred to the Owner’s operations team. The Owner’s team will consist of plant operators and maintenance staff who will enlist the help of suppliers, contractors, engineers and construction management personnel, as needed, to dry run and then wet run the systems until they are finally accepted by the Owner’s operations management. The transfer of systems will be formally documented and will include all mechanical/electrical testing documents and supplier information.

24.13

Mechanical Completion

Mechanical Completion is a term used with contractors, and often defined in contracts to designate the point at which the contractor is considered to have completed his work such that the Owner may operate the facility in a safe manner. The facility may not be completely finished at such time; however, mechanical completion may pertain to a building or a system, for example the fresh water system. Mechanical Completion is often descriptive of Substantial Completion at which time a full punch list of deficiencies remaining is developed by the contractor, the Construction Management team and the Owner, and used to measure the progress to Final Completion. This period in the Project represents the greatest safety risk as the push to complete construction interfaces with the first energization of facilities and equipment. However, before this point is reached, procedures will be established such that electrical lock-out is ensured, hazards identified, and communication protocols established.

24.14

Risk Management

The formal risk management program began during the Feasibility Study phase, and will continue through to commissioning. The project team will review all aspects of the Project throughout the developmental stage, inclusive of environmental, technical, health and safety, community, business and project delivery issues. These reviews will identify the relevant risks and or

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opportunities associated with this Project, assess those risks and opportunities against the outcome objectives and determine the best way to eliminate or control those risks or take advantage of opportunities that may present themselves.

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25.

INTERPRETATION AND CONCLUSIONS

This Feasibility Study indicates that the Rainy River Gold Project, based on the calculated Proven and Probable reserves of 116.3 Mt grading 1.08 g/t Au and 2.76 g/t Ag, can support a 20,000 tpd open pit and a 1,000 tpd underground mine.

25.1

Sampling Method, Approach and Analyses

It is SRK’s opinion that Rainy River Resources used industry best practices to collect, handle and assay core samples collected during the 2005 to 2013 period. All drilling and sampling was conducted by appropriately qualified personnel under the direct supervision of appropriately qualified geologists.

Rainy River has partly relied on the internal quality control measures of the accredited laboratory; however, they have also implemented external analytical quality control measures, consisting of inserting control samples (blanks and certified reference material, and field duplicates) with each batch of core drilling samples submitted for assaying.

In the opinion of SRK, the field sampling and assaying procedures used by Rainy River meet industry best practices.

25.2

Data Verification

It is SRK’s opinion that gold grades can be reasonably reproduced, suggesting that the assay results reported by the primary assay laboratories are sufficiently reliable for the resource estimation used in this Feasibility Study.

On completion of the validation procedures, SRK concludes that the digital database for the Rainy River Gold Project is also reliable for resource estimation.

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25.3

Mineral Resources

SRK reviewed and audited the exploration data available for the Rainy River Gold Project as well as the exploration methodologies adopted to generate this data. Exploration work is professionally managed and procedures are adopted that generally meet accepted industry best practices. SRK is of the opinion that the exploration data are sufficiently reliable to interpret with confidence the boundaries of the gold-rich sulphide mineralization and support evaluation and classification of mineral resources in accordance with generally accepted CIM Estimation of Mineral Resource and Mineral Reserve Best Practices Guidelines and CIM Definition Standards for Mineral Resources and Mineral Reserves.

The drilling database includes information from 1,435 core boreholes (662,849 m), 237 of which (95,760 m) have been drilled in 2012 since the February 2012 mineral resource model. Recent drilling has been mainly infill, which has upgraded the resource classification of the resource model. The mineral resource statement effective October 10, 2012 is tabulated in Table 25-1.

A comparison between the February 2012 and the October 2012 Mineral Resource Statements is shown in Table 25-2. The proportion of Measured and Indicated has increased significantly at the expense of Inferred mineral resources.

SRK considers that the mineral resource model document herein is sufficiently reliable to support engineering and design studies to evaluate the viability of a mining project at a feasibility level.

The newly discovered Intrepid Zone has not been included in any calculations for mining, plant recoveries or financials and does not affect the content of this Report.

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Table 25-1: Mineral Resource Statement1, Rainy River Gold Project, Ontario,

SRK Consulting, October 10, 2012


			Grade				Metal
Category	Quantity ‘000 t		Au g/t		Ag g/t		Au ‘000 oz.		Ag ‘000 oz.
In Pit Mineral Resources²(cut-off grade: 0.35 g/t gold)
Measured		27,550		1.32		1.90		1,168		1,681
Indicated		112,271		1.11		2.51		4,012		9,048
Measured and Indicated		139,821		1.15		2.39		5,180		10,728
Inferred		19,353		0.88		1.40		550		870
Out of Pit Mineral Resources²
Indicated		14,466		0.80		3.84		373		1,785
Inferred		73,555		0.68		2.53		1,610		5,980
Underground Mineral Resources²(cut-off grade: 2.5 g/t gold)
Measured		88		4.97		2.76		14		8
Indicated		4,148		4.50		6.12		600		816
Measured and Indicated		4,236		4.50		6.05		614		824
Inferred		897		4.18		4.63		120		134
Combined Mineral Resources: In Pit, Out of Pit and Underground²
Measured		27,638		1.33		1.90		1,182		1,689
Indicated		130,885		1.18		2.77		4,985		11,649
Measured and Indicated		158,523		1.21		2.62		6,167		13,338
Inferred		93,805		0.75		2.32		2,280		6,984

Mineral resources are reported in relation to conceptual pit shells. On average, the open pit extends to an elevation of 500 m below surface. Mineral resources are not mineral reserves and do not have a demonstrated economic viability. All figures are rounded to reflect the relative accuracy of the estimate. Figures may not add due to rounding. All composites have been capped where appropriate. Qualified Persons - The mineral resource statement was prepared by Dorota El-Rassi, P.Eng. (APEO #100012348) and Glen Cole, P.Geo. (APGO #1416), of SRK, both “independent qualified persons” as that term is defined in National Instrument 43-101. Rainy River’s exploration program in Richardson Township is being supervised by Kerry Sparkes, P.Geo. (APEGBC #25261), Vice-President, Exploration and a Qualified Person as defined by National Instrument 43-101. The Company continues to implement a rigorous QA/QC program to ensure best practices in sampling and analysis of drill core. The estimates of mineral resources may be materially affected by environmental, permitting, legal, title, taxation, sociopolitical, marketing, and other relevant issues.

Open pit mineral resources are reported at a cut-off grade of 0.35 g/t gold, underground mineral resources are reported at a cut-off grade of 2.5 g/t gold based on a gold price of USD $1,100 per ounce, a silver price of USD $22.50 per ounce and a foreign exchange rate of CAD $1.10 to USD $1.00. Metallurgical recoveries include 88% for gold in the open pit mineral resources and 90% for gold in the underground mineral resources with a silver recovery of 75%.

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A comparison between the February 2012 and the October 2012 Mineral Resource Statements is shown in Table 25-2. The additional 237 boreholes drilled in 2012 have resulted in an overall increase in the contained gold metal. It should be noted that a significant portion of this additional mineral resource is classified as Indicated. The proportion of Measured and Indicated has increased significantly at the expense of Inferred mineral resources.

Table 25-2: Comparison of February 2012 and October 2012 Mineral Resource Statements


	Quantity			Grade (g/t)						Contained Metal (oz.)
Classification	(tonnes)			Gold			Silver			Gold			Silver
Open Pit
Measured		19	%		2	%		-5	%		5	%		7	%
Indicated		2	%		3	%		13	%		3	%		9	%
Measured & Indicated		5	%		3	%		10	%		6	%		10	%
Inferred		7	%		-3	%		13	%		5	%		6	%
Underground
Measured		-1	%		7	%		8	%		6	%		7	%
Indicated		40	%		4	%		11	%		24	%		37	%
Measured & Indicated		39	%		4	%		11	%		19	%		37	%
Inferred		-27	%		1	%		-34	%		-28	%		-57	%

25.4

Sampling Preparation, Analysis and Security

In the opinion of SRK, Rainy River personnel used care in the collection and management of field and assaying exploration data. In addition, the sampling preparation, security and analytical procedures used by Rainy River are consistent with generally accepted industry best practices and are therefore adequate.

25.5

Mining Methods and Reserves

The combined open pit and underground mine mineral reserves are summarized in Table 25-3. This summary is reported at two (2) gold equivalent cut-off grades. Open pit mineral reserves are reported at a cut-off of 0.30 g/t Au eq., whereas underground quantities are reported at a cut-off grade of 3.5 g/t Au eq. The summary of mineral reserves includes both dilution and mining recovery factors.

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Table 25-3: Proven and Probable Mineral Reserves (Effective date April 10, 2013)^1,2,3


Reserves Category	Tonnage (Mt)		Au Grade (g/t)		Ag Grade (g/t)		Au (In-Situ oz.)		Ag (In-Situ oz.)
Open Pit
Proven		27.7		1.14		1.94		1,014,584		1,727,979
Probable		85.5		0.91		2.88		2,510,641		7,918,793

TOTAL		113.2		0.97		2.65		3,525,225		9,646,772

Underground
Proven
Probable		3.1		5.07		6.69		506,283		668,240

TOTAL		3.1		5.07		6.69		506,283		668,240


Combined
Proven		27.7		1.14		1.94		1,014,584		1,727,979
Probable		88.6		1.06		3.01		3,016,924		8,587,034

TOTAL		116.3		1.08		2.76		4,031,508		10,315,013

Open pit reserves have been estimated using a cut-off grade of 0.30 g/t gold-equivalent, and underground reserves have been estimated using a cut-off grade of 3.5 g/t gold-equivalent. Open pit reserves have been estimated using a dilution of 9.7% at 0.22 g/t Au and 1.31 g/t Ag, and underground reserves have been estimated using a CAF dilution of 9% at 0.61 g/t Au and 4.16 g/t Ag and LH dilution of 10% at 1.56 g/t Au and 1.28 got Ag. Open pit reserves have been estimated using a mine recovery of 95%, and underground reserves have been estimated using a mine recovery of 95%.

Qualified persons - The mineral reserve statement was prepared by Patrice Live (OIQ #38991) of BBA and Donald Tolfree (APEGBC #32557), of Golder, both “independent qualified persons” as that term is defined in National Instrument 43-101. Rainy River’s engineering assessment in Richardson Township is being supervised by Garett Macdonald, P.Eng. (PEO #90475344), Vice-President, Operations and a Qualified Person as defined by National Instrument 43-101. The estimates of mineral resources may be materially affected by environmental, permitting, legal, title, taxation, sociopolitical, marketing, and other relevant issues.

25.6

Metallurgy and Processing

The results from the SGS testwork program are the basis for the mineral reserve estimate and Feasibility Study. Based on a trade-off study, it was determined that the whole rock leaching option with gravity separation was the most economic alternative and was therefore used as the basis for the Feasibility Study. The main reasons for this selection were the significant amount of energy associated with regrinding the flotation concentrate and the high cyanide consumption in the

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flotation concentrate leaching, in addition to risk associated with ultrafine grinding of this material. All subsequent testwork was based on cyanide leaching of the gravity tailings.

An extensive grinding testwork campaign has allowed for definition of the overall hardness of each zone and indicated that there are several portions of the deposit that will have high energy requirements and this is reflected in the design of the process plant. The design Axb value is 24.2, which corresponds to the 80^th percentile of the testwork data, weighted by zone. The strong correlation between the four (4) methods used to size the grind circuit provides a good level of confidence in the sizing of the SAG and ball mill. The grind size chosen for this study was 75 µm based on a cost versus revenue study performed by BBA.

The process circuit is designed for a throughput of 21,000 tpd and will incorporate primary crushing, semi-autogenous milling with pebble crushing of the oversized material, ball milling, gravity and cyanide leach, followed by gold and silver recovery by carbon-in-pulp (CIP), cyanide destruction system, stripping and electrowinning of the pregnant strip solution.

The process is expected to yield an overall gold recovery including solution losses of approximately 0.3%, ranging from 90-91% (LOM: 90.4%), and a silver recovery including solution losses of approximately 2-3%, ranging from 63-65% over the life of mine (LOM: 64.1%). The gold recovery varies throughout the deposit and the CAP Zone has lower gold recoveries than the other zones. The CAP Zone material mined in the first ten (10) years is stockpiled and processed along with the low grade stockpile material in Years 11 to 16.

25.7

Infrastructure

Buildings

The main administration building will be located at the entrance of the mine site and will house administration and safety/security staff only.

The mine garage will have a total of six (6) maintenance bays, including two (2) bays for auxiliary vehicles and one (1) bay dedicated for welding. The truck wash facility will include a tire change

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bay, designed to allow tire changes from inside the building on only one side of the vehicle. The other side will be accessible from outside the building through two (2) garage doors.

The mine office will be located next to the truck shop and will house the mine, maintenance and mine engineering office staff. The process plant office will be located on the west side of the process building between the leach tanks and the pre-leach thickener and will be connected to the main building via a short corridor.

Roads

Mine haul roads will be built at the start of the Project and will remain in use for the duration of the mine life. Site access roads to the tailings management area and to the explosives plant are already existing roads which will be enlarged and resurfaced with crushed stone.

Based on the finding of the TBT study, the preferred alignment for rerouting Highway 600 around the proposed development area optimizes the use of existing road easements and is the preference of both the Township of Chapple and Rainy River Resources. Access to Marr Road will be provided via Korpi Road and the new East Access Road.

Tailings Management and Dam Design

The total volume of tailings produced over the mine life will be approximately 82 Mm³ at a deposited dry density of 1.4 t/m³, and its location to the northwest of the open pit was selected in consideration of the topography, location of the pit and watershed boundaries, availability of dam construction materials and suitability for a flooded water cover for closure.

The water management system developed by AMEC is designed to: generate a reliable water source for process plant operations and ancillary uses while optimizing the quantity and quality of site effluents released to the environment. Water will be recycled from various manmade ponds for process water, in order to minimize the volume of fresh water to be taken from local watercourses. The system has been designed to ensure a reliable water supply at all times of the year and to allow for contingencies, such as dry years.

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Electricity

The total power demand of the Project was determined to be approximately 56.9 MW based on the estimated connected load, running load and running power. Electricity will be supplied by a proposed new 17 km long 230 kV power line to be built and subsequently connected to the existing 230 kV Hydro One line connecting Fort Frances and Kenora.

25.8

Environmental Permitting

The process of environmental permitting is well understood and a preliminary schedule outlining the critical steps has been developed in this study and has been integrated into the preliminary Project execution schedule. The obtaining of environmental approvals is on the Project’s critical path and no construction activities can commence until the required permits and authorizations are obtained.

There is considerable environmental baseline information currently available regarding the site and the surrounding area, compiled through extensive field investigations conducted over a 5-year period. Based on the information available to-date, there are no environmental aspects that are considered to be limiting to the Project’s development, as the Project design has considered appropriate environmental mitigation measures including avoidance of critical areas as practical.

25.9

Financial Analysis

This Feasibility Study indicates that the Rainy River Gold Project, based on the Proven and Probable Mineral Reserves (Open Pit and Underground reserve quantities) of 116.3 Mt grading 1.08 g/t Au and 2.76 g/t Ag, can support an 20,000 tpd open pit and a 1,000 tpd underground mine. Mineralized material will be sent to a process plant designed to achieve a gold and silver recovery of approximately 90.4% and 64.1% over the life-of-mine, respectively. It is anticipated that, over a mine life of 16 years, approximately 3,645 koz. of gold and 6,615 koz. of silver will be produced.

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The initial capital cost of the Project is estimated to be $713M; with open pit sustaining capital of $322M and underground development phase and sustaining capital of $68M and $95M, respectively. The life-of-mine cash operating cost is USD $544/oz. Au, including silver credits and royalty payments. The first ten (10) years of operation will have a cash operating cost of USD $468/oz. Au. The pre-tax Project NPV is estimated to be $1,296M using a discount rate of 5%. The Project’s internal rate of return (IRR) pre-tax is estimated to be 27.8% and the simple payback period of the Project is 3.1 years.

The Rainy River Gold Project is subject to three (3) levels of taxation, including federal income tax, provincial income tax and provincial mining taxes. Rainy River compiled the taxation calculations for the Project with assistance from third-party taxation experts, and the calculation includes the impact of changes announced in the March 2013 federal budget. The three (3) taxes result in a combined effective tax rate of 28.5% on the Project.

After tax NPV is estimated to be $931M using a discount rate of 5%. The Project IRR (after tax) is estimated to be 23.7% and the simple after tax payback period is 3.2 years.

The after-tax sensitivity analysis indicates that positive project returns can be achieved over the likely range of variation in gold prices (± 20%), metal recovery (- 10% / +3%), capital costs (± 20%) and operating costs (± 20%). The project financials are most sensitive to metal price and metal recovery.

25.10

Conclusion

Based on the information available and degree of development of the Project as of the effective date of this Report, it is BBA’s opinion that the Rainy River Gold Project is sufficiently robust, both technically and financially, to warrant proceeding to the next stage of project development, consisting of final design and Detailed Engineering. This conclusion is based on the recommendations and work plan as presented in Section 26.

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26.

RECOMMENDATIONS

The following recommendations are made considering the results of this Feasibility Study and the project risks identified. BBA recommends a work program that includes further exploration drilling, metallurgical testing and various studies aiming at completing the characterization of the Project in preparation for the Detailed Engineering Study phase. The suggested work program includes the following components:

•

Continuation of open pit and underground mine design optimization;

•

Procurement of long lead time mining and process equipment;

•

Continuation of preparatory work to secure electrical power and procurement of long lead time electrical equipment;

•

Continue recruiting key personnel;

•

Continuation of environmental assessment studies;

•

Continuation of First Nation and public consultations;

•

Continuation of hydrogeological studies in specific areas;

•

Secure Project financing;

•

Secure required permits and authorizations from government and regulatory agencies;

•

Initiate detailed engineering activities; and

•

Construction of the project (following appropriate approvals).

26.1

Proposed 2013 Work Program

The costs of this next engineering phase and additional exploration activities are estimated at approximately CAD $39.6M, as shown in Table 26-1. As of the effective date of this report, Rainy River has already authorized and initiated a considerable amount of the recommended work.

Table 26-1: Budget for 2013


Activity	Cost ($ M)
Exploration		18.9
Condemnation Drilling		2.7
Basic and Detailed Engineering Activities		18.0

TOTAL		39.6

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BBA has reviewed Rainy River Resources’ proposal for further exploration and studies on the Rainy River Gold Project property and considers that the budget for the proposed program is reasonable. BBA recommends that Rainy River Resources implement the program as proposed, subject to either funding or other matters that may cause the proposed program to be altered in the normal course of its business activities, or alterations which may affect the program as a result of exploration activities themselves.

26.2

Further Recommendations

During the completion of the Feasibility Study, a large amount of work has been completed to bring the various aspects of the Project to its current stage (May 2013). Further development of the Rainy River Gold Project property should address key risks and opportunities that have been identified through the feasibility work program.

Geology and Exploration:

•

Additional drilling is required to infill the remnant gaps in the drilling data with the potential to increase the mineral resources; infill areas of inferred resources to improve resource classification; and test the lateral and depth extensions of the gold mineralization;

•

Geological studies aimed at improving the understanding of the geological and structural setting of the deposit particularly within the silver-enriched zones and the mafic-hosted nickel-copper rich zones, followed by revised 3D modeling;

•

Further definition of drilling and geological modeling of gold and silver mineralization within other zones in close proximity to currently reported resources, such as the Intrepid Gold-Silver Zone; and

•

Condemnation drilling to support mine infrastructure design.

Rainy River has a condemnation drilling program designed for 2013 comprising of 20,000 m of core drilling at a budgeted cost of about CAD $2.7M. In addition, Rainy River is planning an additional 70,000 m of core drilling throughout 2013 on mineral resource targets at a budgeted cost of about CAD $18.9M. This proposed condemnation/exploration drilling program is considered to be aligned with the recommendations of this Study.

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Open Pit and Underground Mining

The mining aspects of the Feasibility Study, including the mine design, the mine planning optimization process, mining dilution and fleet selection have been carried out according to industry standards and all risks and concerns have been identified and addressed. BBA recommends the following work:

•

To review the current overburden material stockpiling approach;

•

Revisit the elevated cut-off grade strategy using the current mill throughput and mining method to maximize the positive NPV impact;

•

To generate a 10 m x 10 m x 10 m block model for the open pit operation to match the size of the selected equipment;

•

To review the pit slope as more data is collected and available;

•

To develop a detailed operational mine plan for pre-production and for the first two (2) years of operation;

•

Underground definition drilling to improve the understanding of the factors controlling the grade;

•

Optimizing the cut and fill design possibly using a smaller resource model block size;

•

Update ventilation design;

•

Update underground hydrogeological estimates and dewatering design to account for potential pit inflows underground; and

•

Complete cemented rock fill (“CRF”) testing to confirm the backfill strengths and mix design.

Open Pit Geotechnical Design

At the next stage of detailed engineering refinement in the pit design will require the modification of bench face angles primarily in the identified zones of 3 t and 7 t. This will require the bench face angle to transition in the 3 t region from 55 degrees to 70 degrees from west to east, and in the 7 t region also transition from 55 degrees to 70 degrees from north to south.

The design specifications for Zone 3, the north wall of the ODM/17 are primarily based on the geotechnical drilling in Zone 1 (south wall) of the ODM/17, with confirmation from the televiewer data of the exploration boreholes in the central core of the pit, and are adequate for this level of study. It is recommended however to confirm variation in the dip of the controlling foliation set from

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east to west and north to south with either additional televiewer surveys or two (2) additional boreholes either side of the pit centre towards the north.

Underground Geotechnical Design

•

The main underground zones of the ODM/17 have been investigated with three (3) geotechnical boreholes and supplemented with televiewer surveys from exploration boreholes in the central zone. The level of information is adequate for the present Study, however, it is recommended that further investigations are performed in line with other geotechnical bedrock drilling;

•

All further exploration core should be photographed prior to splitting and logged to record rock quality designation (“RQD”) values. This information is significantly useful to supplement any directed geotechnical program;

•

The underground water inflow estimates and sump design should be updated in the next phase of work to consider inflow from the open pit into the stopes.

•

Geotechnical investigation should be performed for the portal location and for the first 300 m of the ramp;

•

Geotechnical investigation should be performed for all major raises from surface to the underground mine. This should be in the form of geotechnical pilot holes for the main ventilation shafts;

•

Variance in the open pit and underground mining geometry due to changes in the cut-off grade should be investigated as to their impact on infrastructure through rock mechanics and hydrogeological studies;

•

Long-term backfill testing should be completed; and

•

It is recommended that once the ramp has reached a depth approaching 500 m, that in-situ stress overcoring tests are performed preferably at two locations separated vertically by 100 m to 150 m.

Metallurgical Testing and Processing:

•

Hydraulic testing for the final confirmation of tailings pumping characteristics and thickener settling characteristics;

•

Investigate the design circulating load for the ball mill circuit to potentially improve grinding efficiency;

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•

The height difference between the leach tanks can be optimized based on the viscosity of the material. A conservative number of 0.7 m height difference between each tank was used for the Feasibility Study;

•

Additional testwork or modelling of the electrowinning cells should be done to confirm the sizing is sufficient given the high silver content of the deposit; and

•

Review design of the trash screens and safety screen to ensure ease of operation.

Infrastructure

•

The site road designs and aggregate usage assumptions are recommended to be revisited prior to construction start-up;

•

Additional drilling to obtain the rock profile at the primary crusher is required in order to finalize the mechanically stabilized earth wall design; and

•

Submit application for electrical power to IESO Ontario.

Environmental and Tailings Management:

•

Environmental assessment(s) and consultation/engagement activities should proceed with the objective of gaining environmental approval for the Project in line with the overall Project schedule;

•

Geochemical characterization should continue as the processing and mining plans are detailed, with modification to the mineral waste management plan as appropriate;

•

Estimated reclamation costs and bonding requirements should be reassessed in the next phase of development as more detailed engineering designs become available;

•

Supplemental geotechnical site investigations at the dam sites are required to better define the foundation conditions for construction material quantity estimation, particularly for the west dam of the TMA. Shallow or exposed bedrock foundations may require treatment based on the rock quality; and

•

Additional work should be carried out for delineation and characterization of the mine waste overburden suitable for dam construction. A clay excavation and berm construction trial is recommended to observe the excavation, handling and trafficability characteristics of the clays.

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27.

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SRK Consulting (Canada) Inc., 2011. Mineral Resource Evaluation, Rainy River Gold Project, Western Ontario, Canada prepared for Rainy River Resources Ltd. Public document filed on SEDAR, 133 pages, dated April 8, 2011.

SRK Consulting (Canada) Inc., 2011. Mineral Resource Evaluation, Rainy River Gold Project, Western Ontario, Canada prepared for Rainy River Resources Ltd., 56 pages, dated August 11, 2011.

SRK Consulting (Canada) Inc., 2012. Mineral Resource Evaluation, Rainy River Gold Project, Western Ontario, Canada prepared for Rainy River Resources Ltd., 299 pages, dated April 9, 2012.

Stanley, J.E. 2012. Layout & Design Equipment Budget Estimate Pre-Engineering Specifications. Report compiled and presented to Paolo A. Toscano (Rainy River Resources), Director of Metallurgy. June 2012.

Stanley, J.E. 2013. Layout & Design Equipment Budget Estimate Pre-Engineering Specifications (25 K Plant), Report compiled and presented to Paolo A. Toscano (Rainy River Resources), Director of Metallurgy. February 2013.

Starkey & Associates, 2012. SAGDesign Test Work Results for 7 Samples. Test Work Validation Report prepared for BBA and Rainy River Resources Ltd, dated June 11, 2012.

Statistics Canada, Capital Expenditure Price Statistics (Volume 26, N°. 4). April 2010

TBT Engineering Consulting Group, 2012. Highway 600 Realignment. Feasibility Study prepared for Rainy River Resources Ltd., dated February 9, 2012.

Wartman, Jakob, M., 2011 – Physical Volcanology and Hydrothermal Alteration of the Rainy River Gold Project, northwest Ontario. 154 pages.

http://www.d.umn.edu/geology/research/thesis.html.

Whittle Consulting Pty Ltd. (Australia), 2012. Enterprise Optimization Rainy River Resources Ltd Rainy River Gold Project. Prepared by Richard Peevers. Reviewed by Gerald Whittle. July 2012.

27-10

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

APPENDIX A

Rainy River Resources Title Opinion

(as of March 13, 2013)

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Of Counsel

The Right Honourable Pierre Elliott Trudeau, P.C., C.C., C.H., Q.C., FRSC (1984 - 2000)

The Right Honourable Jean Chrétien, P.C., C.C., O.M., Q.C.

The Honourable Donald J. Johnston, P.C., O.C., Q.C.

Pierre Marc Johnson, G.O.Q., FRSC

The Honourable Michel Bastarache, C.C.

The Honourable René Dussault, FRSC

The Honourable John W. Morden

Peter M. Blaikie, Q.C.

André Bureau, O.C.

March 13, 2013

Rainy River Resources Ltd.

1 Richmond Street West, Suite 701

Toronto, ON M5H 3W4

Dear Sirs/Mesdames:

Re:

Confirmation of title to or other interest in land holdings related to NI 43-101 Filing by Rainy River Resources Ltd.

Our Reference: 063125-0011

We have acted as counsel to Rainy River Resources Ltd. (“Rainy River”) for the purpose of providing this opinion in connection with the above-noted matter.

Scope of Examination and Searches

For the purpose of providing this opinion, we have examined:

(a)

registered title to the patented lands described in Schedule “A” to this opinion (the “Patented Lands”) for the sole purpose of noting the registered owner noted thereon, as disclosed by the relevant records (the “Registers”) of the Land Registry Office for the Land Titles Division of Rainy River (No. 48) (the “LRO”);

(b)	registered title to the patented leasehold lands described in Schedule “B” to this opinion (the “Leasehold Lands”) for the sole purpose of noting the registered leasehold owner noted thereon, as disclosed by the relevant Registers;

(c)	the relevant Registers for the patented lands described in Schedule “C” to this opinion (the “Option Lands”) for the sole purpose of noting whether notice of an option to purchase in favour of Rainy River is noted thereon;

(d)

copies of the active mining claims abstract summaries and transactions listings current to March 13, 2013 maintained by the Mining Recorder’s Office of the Ontario Ministry of Northern Development and Mines (the “Records”) for each of the unpatented claims described in Schedule “D” (the “Unpatented Claims”) and for the unpatented claims described in the English Option Agreements, the Roisin Option Agreement and the Timberridge Land & Forestry Services Inc. and described on

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Schedule “D1” (the “Optioned Unpatented Claims”) for the sole purpose of noting the recorded holder thereof;

(e)	photostatic copies of two purchase option agreements dated March 3, 2010 between Perry English for Rubicon Minerals Corporation, as optionor and Rainy River, as optionee (collectively, the “English Option Agreements”) relating to the applicable Optioned Unpatented Claims,

(f)	photostatic copy of an option agreement dated December 16, 2011 between Fred A. Roisin, as optionor and Rainy River, as optionee (collectively, the “Roisin Option Agreement”) relating to the applicable Optioned Unpatented Claims,

(g)	photostatic copy of an option agreement dated March 29, 2012 between Timberridge Land & Forestry Services Inc., as optionor and Rainy River, as optionee (collectively, the “Timberridge Option Agreement”) relating to the applicable Optioned Unpatented Claims,

and our opinions expressed below are expressly subject and limited to the results of such examinations. With your concurrence, we have not conducted any other or more in depth reviews, examinations or searches, including of the relevant Registers or the Records, other than those indicated above and our opinion is therefore qualified as to any matters that would be revealed by any such other or more in depth reviews, examinations or searches.

Assumptions and Reliances

In connection with the opinions expressed in this letter, we have assumed:

(a)

in the conduct of our examination above, the genuineness of all signatures, the legal capacity of all individuals who have executed documents or instruments, the authenticity of all documents and instruments submitted to or reviewed by us as originals, the conformity to originals of all documents and instruments submitted to or reviewed by us as certified, telecopied or photostatic copies or facsimiles thereof, and the authenticity of the originals of such certified copies, photocopies, telecopies or facsimiles, and the enforceability of all such original, certified, telecopied, photostatic or facsimile copies of such documents and instruments;

(b)

that any persons purporting to have executed the documents or instruments examined in the course of our examinations noted above, are, in fact, the same persons named therein, and, when executed by a corporation or governmental authority, that the persons so executing on behalf of the corporation or governmental authority have been duly and validly authorized to do so as signing officers of the corporation or governmental authority;

(c)

any previous or current registered or recorded corporate owners of the Patented Lands, Leasehold Lands, Unpatented Claims or Optioned Unpatented Claims, or of an interest in and to the Patented Lands, Leasehold Lands or Unpatented Claims or Optioned Unpatented Claims, were duly incorporated and validly existing in their jurisdiction of incorporation, were extra-provincially registered in the Province of Ontario and were entitled to own, and had the corporate power and capacity to own, property or an interest in property in the Province of Ontario and to execute and deliver all agreements or instruments, and that such corporations were not dissolved or wound up, voluntarily or involuntarily, during the period that such corporations were the registered or recorded owners or holders of the Patented Lands, Leasehold Lands, Unpatented Claims or Optioned

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Unpatented Claims or of an interest in and to the Patented Lands, Leasehold Lands, Unpatented Claims or Optioned Unpatented Claims;

(d)	the accuracy and currency of the indices and filing systems maintained at any public offices where we have conducted reviews, examinations or searches or made enquiries or caused such reviews, examinations or searches or enquiries to be conducted or made;

(e)

that the Registers and Records pertaining to title to the Patented Lands, Leasehold Lands, Option Lands, Unpatented Claims and Optioned Unpatented Claims are accurate and that there are no unrecorded transfers, assignments, agreements or other unrecorded encumbrances or rights, title or interests affecting such lands or the registered or recorded owners’ or holders’ interest therein;

(f)

each party to the English Option Agreements, the Roisin Option Agreement and the Timberridge Option Agreement (collectively, the “Option Agreements”), including Rainy River, is a validly created and subsisting legal entity, has all necessary power and capacity to execute, deliver and perform the Option Agreements, has duly authorized, executed and delivered the Option Agreements and such Option Agreements constitute legal, valid and binding obligations of such parties, enforceable against them in accordance with their terms; and

(g)	the Option Agreements and the underlying leases for the Leasehold Lands are in good standing and in full force and effect as of the dates noted in paragraph 2 below, have not been cancelled or terminated or assigned, and all parties to such agreements have performed all obligations thereunder.

We are solicitors qualified to carry on the practice of law in the Province of Ontario. The opinion expressed herein extends only to the laws of the Province of Ontario and the federal laws of Canada applicable therein in force as of the date of this opinion.

Opinion

Based upon and subject to the foregoing, and subject to the exceptions and qualifications to title set out in Schedule “E”, we are of the opinion that:

1.	as of March 13, 2013, Rainy River is noted on the relevant Registers as the registered owner in fee simple of each of the Patented Lands described on Schedule “A”;

2.	as of March 13, 2013, Rainy River is noted on the relevant Registers as the registered holder of a leasehold estate for each of the Leasehold Lands described on Schedule “B”;

3.	as of March 13, 2013, certain notices of an option to purchase in favour of Rainy River have been registered against the relevant Registers for each of the Option Lands described on Schedule “C”;

4.	as of March 13, 2013, Rainy River is noted on the relevant Records as the recorded holder of the Unpatented Claims described in Schedule “D”; and

as of March 13, 2013, the Option Agreements create a contractual interest in the nature of an option to acquire a 100% interest in favour of Rainy River in and to the Optioned Unpatented Claims described in Schedule “D1”. We note that as of such date, the Option Agreements are not recorded against the relevant Records for the Optioned Unpatented Claims.

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This opinion is provided solely for the internal use of the addressees hereof and may not be quoted or otherwise referred to in other documents or relied upon by either of you for any purpose or in conjunction with any matter or transaction, other than the matter noted above, nor may it be quoted or used or relied on by any other person in conjunction with any matter or transaction, without our express written permission.

Yours truly,

/s/ Heenan Blaikie

HEENAN BLAIKIE LLP

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Schedule “A”

Patented Lands

1.	56036-0084 (LT)

PCL 17371 SEC RAINY RIVER; N 1/2 LT 9 CON 6 MATHER EXCEPT PT COVERED BY WATERS OF PINE RIVER; CHAPPLE

2.	56036-0077 (LT)

PCL 16847 SEC RAINY RIVER; N 1/2 LT 12 CON 6 MATHER EXCEPT PT 2, 48R1197; CHAPPLE

3.	56035-0066 (LT)

PCL 16077 SEC RAINY RIVER; N 1/2 LT 12 CON 1 POTTS; CHAPPLE

4.	56035-0098 (LT)

PCL 18689 SEC RAINY RIVER; S 1/2 LT 12 CON 3 POTTS; CHAPPLE

5.	56035-0176 (LT)

PCL 5899 SEC RAINY RIVER; S 1/2 LT 12 CON 1 POTTS; CHAPPLE

6.	56035-0090 (LT)

PCL 17941 SEC RAINY RIVER; S 1/2 LT 12 CON 2 POTTS; CHAPPLE

7.	56035-0242 (LT)

SURFACE RIGHTS ONLY; N 1/2 LT 11 CON 2 POTTS; CHAPPLE

8.	56042-0082 (LT)

PCL 22495 SEC RAINY RIVER; E 1/2 OF N 1/2 LT 7 CON 2 RICHARDSON EXCEPT A55914 BEING THE MINING RIGHTS; CHAPPLE

9.	56042-0189 (LT)

MINING RIGHTS ONLY, E 1/2 OF S 1/2 LT 9 CON 2 RICHARDSON EXCEPT PT 5 48R1985 & PL S-439; CHAPPLE

10.	56042-0187 (LT)

MINING RIGHTS ONLY, E 1/2 OF N 1/2 LT 9 CON 1 RICHARDSON EXCEPT PT 6 PL S-439, PT 4 48R1985; CHAPPLE

11.	56042-0160 (LT)

SURFACE RIGHTS ONLY, S 1/2 LT 4 CON 2 RICHARDSON EXCEPT PL S-391, SLT30449; CHAPPLE

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12.	56042-0161 (LT)

MINING RIGHTS ONLY, S 1/2 LT 4 CON 2 RICHARDSON EXCEPT PL S-391, SLT30449; CHAPPLE

13.	56042-0162(LT)

SURFACE RIGHTS ONLY, N 1/2 LT 4 CON 3 RICHARDSON; CHAPPLE

14.	56042-0163 (LT)

MINING RIGHTS ONLY, N 1/2 LT 4 CON 3 RICHARDSON; CHAPPLE

15.	56042-0097 (LT)

PCL 25891 SEC RAINY RIVER; PT LT 6 CON 1 RICHARDSON AS IN A68613; MINING RIGHTS ONLY; CHAPPLE

16.	56042-0034 (LT)

PCL 14408 SEC RAINY RIVER; PT LT 6 CON 1 RICHARDSON AS IN SLT53957 EXCEPT PL S-391, PT 1 48R1961 & A68613 BEING THE MINING RIGHTS; CHAPPLE

17.	56042-0173 (LT)

MINING RIGHTS ONLY; S 1/2 LT 9 CON 3 RICHARDSON; CHAPPLE

18.	56042-0139 (LT)

PCL 26007 SEC RAINY RIVER MRO; N 1/2 LT 7 CON 2 RICHARDSON EXCEPT THE E 1/2; CHAPPLE

19.	56042-0104 (LT)

PCL 4534 SEC RAINY RIVER; N 1/2 LT 7 CON 2 RICHARDSON EXCEPT THEE 1/2 & A73222 BEING THE MINING RIGHTS; CHAPPLE

20.	56042-0102 (LT)

PCL 26007 SEC RAINY RIVER MRO; S 1/2 LT 8 CON 2 RICHARDSON EXCEPT PT 3 PL S-439, THE N 1/2, & PT 1 & 2 48R1985; CHAPPLE

21.	56042-0113 (LT)

PCL 5483 SEC RAINY RIVER; S 1/2 LT 8 CON 2 RICHARDSON EXCEPT PT 3, PL S439, THEN 1/2, PT 1 & 2 48R1985, A73222 BEING THE MINING RIGHTS; CHAPPLE

22.	56042-0179 (LT)

MINING RIGHTS ONLY; W 1/2 OF N 1/2 LT 12 CON 3 RICHARDSON; TOWNSHIP OF CHAPPLE

23.	56042-0177 (LT)

MINING RIGHTS ONLY; PT LT 3 CON 1 RICHARDSON AS IN SLT76289; CHAPPLE

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24.	56042-0100 (LT)

PCL 25984 SEC RAINY RIVER; N 1/2 LT 6 CON 2 RICHARDSON MINING RIGHTS ONLY; CHAPPLE

25.	56042-0128 (LT)

PCL 8825 SEC RAINY RIVER; N 1/2 LT 5 CON 2 RICHARDSON EXCEPT A72814 BEING THE SURFACE RIGHTS; CHAPPLE

26.	56042-0098 (LT)

PCL 25892 SEC RAINY RIVER; S 1/2 LT 5 CON 2 RICHARDSON MINING RIGHTS ONLY; CHAPPLE

27.	56042-0175 (LT)

MINING RIGHTS ONLY; S 1/2 LT 11 CON 2 RICHARDSON EXCEPT THE W 1/2 & PT 11 PL S-439; CHAPPLE

28.	56042-0171 (LT)

MINING RIGHTS ONLY; S 1/2 LT 12 CON 2 RICHARDSON EXCEPT PT 11 PL S446, PT 17, PL S439, SLT50657; CHAPPLE

29.	56042-0169 (LT)

MINING RIGHTS ONLY; N 1/2 LT 12 CON 2 RICHARDSON EXCEPT PT 15 PL S446; TOWNSHIP OF CHAPPLE

30.	56042-0033 (LT)

PCL 14407 SEC RAINY RIVER;PT LT 6 CON 1 RICHARDSON AS IN SLT53956 EXCEPT PL S-391 & A68696 BEING THE MINING RIGHTS; CHAPPLE

31.	56042-0099 (LT)

PCL 25894 SEC RAINY RIVER; FIRSTLY W 1/2 LT 6 CON 1 RICHARDSON SECONDLY E 1/2 LT 6 CON 1 RICHARDSON AS IN SLT53956 MINING RIGHTS ONLY; CHAPPLE

32.	56042-0058 (LT)

PCL 16927 SEC RAINY RIVER; N 1/2 OF S 1/2 LT 6 CON 3 RICHARDSON; CHAPPLE

33.	56042-0166 (LT)

SURFACE RIGHTS ONLY; N 1/2 LT 11 CON 2 RICHARDSON; CHAPPLE

34.	56042-0167 (LT)

MINING RIGHTS ONLY; N 1/2 LT 11 CON 2 RICHARDSON; CHAPPLE

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35.	56042-0158 (LT)

MINING RIGHTS ONLY; N 1/2 LT 4 CON 2 RICHARDSON TOWNSHIP OF CHAPPLE

36.	56042-0159 (LT)

SURFACE RIGHTS ONLY; N 1/2 LT 4 CON 2 RICHARDSON; TOWNSHIP OF CHAPPLE

37.	56042-0116 (LT)

PCL 5939 SEC RAINY RIVER; N 1/2 LT 5 CON 1 RICHARDSON EXCEPT PL S-391, PT 5-7 48R1717 & PT 1 48R1728; CHAPPLE

38.	56042-0114 (LT)

PCL 5614 SEC RAINY RIVER; S 1/2 LT 5 CON 1 RICHARDSON EXCEPT PL S-391 & PT 2 48R1717; CHAPPLE

39.	56042-0063 (LT)

PCL 17725 SEC RAINY RIVER; E 1/2 OF S 1/2 LT 7 CON 3 RICHARDSON; CHAPPLE

40.	56042-0064 (LT)

PCL 17726 SEC RAINY RIVER; S 1/2 LT 3 CON 3 RICHARDSON; CHAPPLE

41.	56042-0060 (LT)

PCL 17110 SEC RAINY RIVER; S 1/2 LT 6 CON 2 RICHARDSON EXCEPT PL S-391; CHAPPLE

42.	56042-0036 (LT)

PCL 14462 SEC RAINY RIVER; PT LT 7 CON 1 RICHARDSON AS IN SLT53727; CHAPPLE

43.	56042-0146 (LT)

MINING RIGHTS IN AND UNDER PT LT 6 CON 3 RICHARDSON AS IN SP3235; CHAPPLE

44.	56042-0147 (LT)

PT LT 6 CON 3 RICHARDSON AS IN SP3235 EXCEPT MINING RIGHTS IN AND UNDER AS IN RD2483; CHAPPLE

45.	56042-0077 (LT)

PCL 21129 SEC RAINY RIVER; PT LT 6 CON 3 RICHARDSON AS IN SLT97369; CHAPPLE

46.	56042-0145 (LT)

PCL 16779 SEC RAINY RIVER; N1/2 OF N1/2 LT 4 CON 1 RICHARDSON EXCEPT PL S391; CHAPPLE

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47.	56042-0053 (LT)

PCL 16630 SEC RAINY RIVER; W 1/2 OF S1/2 LT 9 CON 2 RICHARDSON EXCEPT PT 7, PL S439; CHAPPLE

48.	56042-0164 (LT)

SURFACE RIGHTS ONLY; PT LT 3 CON 1 RICHARDSON AS IN SLT57341; TOWNSHIP OF CHAPPLE

49.	56042-0165 (LT)

MINING RIGHTS ONLY; PT LOT 3 CON 1 RICHARDSON AS IN SLT57341; CHAPPLE

50.	56042-0180 (LT)

SURFACE RIGHTS ONLY; PT LT 3 CON 1 RICHARDSON AS IN SLT30924; CHAPPLE

51.	56042-0181 (LT)

MINING RIGHTS ONLY; PT LT 3 CON 1 RICHARDSON AS IN SLT30924; CHAPPLE

52.	56042-0184 (LT)

SURFACE RIGHTS ONLY; N 1/2 OF S 1/2 LT 4 CON 1 RICHARDSON EXCEPT PL S-391, PT 3 48R1717; CHAPPLE

53.	56042-0185 (LT)

MINING RIGHTS ONLY; N 1/2 OF S 1/2 LT 4 CON 1 RICHARDSON EXCEPT PL S-391, PT 3 48R1717; CHAPPLE

54.	56042-0182 (LT)

SURFACE RIGHTS ONLY; S 1/2 OF S 1/2 LT 4 CON 1 RICHARDSON EXCEPT PL S-391, PT 1-4 48R1191, PT 1 48R1717; CHAPPLE

55.	56042-0183 (LT)

MINING RIGHTS ONLY; S 1/2 OF S 1/2 LT 4 CON 1 RICHARDSON EXCEPT PL S-391, PT 1-4 48R1191, PT 1 48R1717; CHAPPLE

56.	56042-0149 (LT)

MINING RIGHTS ONLY: N PT BROKEN LT 7 CON 1 RICHARDSON BEING ALL THAT PT OF THE SAID LT LYING N OF THE LINE DRAWN ACROSS SAID LT ON A COURSE W ASTRONOMICALLY FROM A POINT ON THE E LIMIT THEREOF. 40 CHAINS S OF THE NW ANGLE OF SAID LT EXCEPT PT 2, PL S439; TOWNSHIP OF CHAPPLE

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57.	56042-0148 (LT)

SURFACE RIGHTS ONLY: N PT BROKEN LT 7 CON 1 RICHARDSON BEING ALL THAT PT OF THE SAID LT LYING N OF THE LINE DRAWN ACROSS SAID LT ON A COURSE W ASTRONOMICALLY FROM A POINT ON THE E LIMIT THEREOF 40 CHAINS S OF THE NW ANGLE OF SAID LT EXCEPT PT 2, PL S439; TOWNSHIP OF CHAPPLE

58.	56042-0047 (LT)

PCL 15881 SEC RAINY RIVER; S 1/2 LT 2 CON 2 RICHARDSON; CHAPPLE

59.	56042-0012 (LT)

PCL 11853 SEC RAINY RIVER; N 1/2 LT 2 CON 1 RICHARDSON; CHAPPLE

60.	56042-0062 (LT)

PCL 17392 SEC RAINY RIVER; W 1/2 OF S 1/2 LT 8 CON 1 RICHARDSON; CHAPPLE

61.	56042-0037 (LT)

PCL 14604 SEC RAINY RIVER; N 1/2 OF S 1/2 LT 4 CON 3 RICHARDSON; CHAPPLE

62.	56042-0129 (LT)

PCL 9080 SEC RAINY RIVER; S 1/2 LT 4 CON 3 RICHARDSON EXCEPT THE N 1/2 ; CHAPPLE

63.	56042-0061 (LT)

PCL 17154 SEC RAINY RIVER; N 1/2 LT 6 CON 2 RICHARDSON EXCEPT A72582 BEING THE MINING RIGHTS; CHAPPLE

64.	56042-0044 (LT)

PCL 15282 SEC RAINY RIVER; S 1/2 OF S 1/2 LT 7 CON 4 RICHARDSON; CHAPPLE

65.	56042-0088 (LT)

PCL 23322 SEC RAINY RIVER; PT LT 6 CON 1 RICHARDSON PT 1 48R1961; CHAPPLE

66.	56042-0065 (LT)

PCL 17752 SEC RAINY RIVER; W 1/2 OF S 1/2 LT 9 CON 1 RICHARDSON RESERVING A STRIP OF LAND TEN FT ALONG THE SHORES OF THE PINE RIVER; CHAPPLE

67.	56042-0101 (LT)

PCL 25991 SEC RAINY RIVER; N 1/2 LT 5 CON 2 RICHARDSON SURFACE RIGHTS ONLY; CHAPPLE

68.	56042-0021 (LT)

PCL 13137 SEC RAINY RIVER; S 1/2 LT 11 CON 1 RICHARDSON; CHAPPLE

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69.	56042-0003 (LT)

PCL 10273 SEC RAINY RIVER; N 1/2 LT 12 CON 1 RICHARDSON EXCEPT PT 16, PL S439; CHAPPLE

70.	56042-0024 (LT)

PCL 13467 SEC RAINY RIVER; N 1/2 LT 11 CON 1 RICHARDSON EXCEPT THE E 1/2 PT 14, PL S439; CHAPPLE

71.	56042-0050 (LT)

PCL 16307 SEC RAINY RIVER; N 1/2 LT 10 CON 3 RICHARDSON; CHAPPLE

72.	56042-0052 (LT)

PCL 16343 SEC RAINY RIVER; W 1/2 OF S 1/2 LT 11 CON 2 RICHARDSON EXCEPT PT 12 & 13, S439; CHAPPLE

73.	56042-0011 (LT)

PCL 11409 SEC RAINY RIVER; S 1/2 LT 5 CON 2 RICHARDSON EXCEPT PL S-391, PT 8 48R1717 & A68624 BEING THE MINING RIGHTS; CHAPPLE

74.	56042-0018 (LT)

PCL 12324 SEC RAINY RIVER; N 1/2 LT 5 CON 3 RICHARDSON; CHAPPLE

75.	56042-0081 (LT)

PCL 22190 SEC RAINY RIVER; S 1/2 LT 5 CON 3 RICHARDSON; CHAPPLE

76.	56042-0038 (LT)

PCL 14665 SEC RAINY RIVER; W 1/2 OF N 1/2 LT 9 CON 1 RICHARDSON EXCEPT PT 8, PL S439; CHAPPLE

77.	56042-0056 (LT)

PCL 16820 SEC RAINY RIVER; S N/2 OF N 1/2 LT 8 CON 2 RICHARDSON; CHAPPLE

78.	56042-0002 (LT)

PCL 10152 SEC RAINY RIVER; N 1/2 LT 11 CON 3 RICHARDSON; CHAPPLE

79.	56042-0055 (LT)

PCL 16754 SEC RAINY RIVER; S 1/2 LT 11 CON 3 RICHARDSON; CHAPPLE

80.	56042-0059 (LT)

PCL 16956 SEC RAINY RIVER; S 1/2 OF N 1/2 LT 8 CON 2 RICHARDSON; CHAPPLE

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81.	56042-0029 (LT)

PCL 14200 SEC RAINY RIVER; S 1/2 LT 12 CON 3 RICHARDSON EXCEPT PT 16, PL S446, PT 2 & 3, PL S455; CHAPPLE

82.	56042-0005 (LT)

PCL 10746 SEC RAINY RIVER; N 1/2 LT 10 CON 2 RICHARDSON; CHAPPLE

83.	56042-0117 (LT)

PCL 6520 SEC RAINY RIVER; N 1/2 LT 9 CON 2 RICHARDSON; CHAPPLE

84.	56042-0112 (LT)

PCL 5455 SEC RAINY RIVER; S 1/2 LT 10 CON 2 RICHARDSON EXCEPT PT 9 PL S439; CHAPPLE

85.	56042-0016 (LT)

PCL 12083 SEC RAINY RIVER; S 1/2 LT 10 CON 1 RICHARDSON; CHAPPLE

86.	56042-0121 (LT)

PCL 7654 SEC RAINY RIVER; N 1/2 LT 10 CON 1 RICHARDSON EXCEPT PT 10, PL S439; CHAPPLE

87.	56042-0133 (LT)

PCL 9665 SEC RAINY RIVER; S 1/2 LT 2 CON 1 RICHARDSON; CHAPPLE

88.	56042-0025 (LT)

PCL 13514 SEC RAINY RIVER; E 1/2 OF N 1/2 2 LT 11 CON 1 RICHARDSON EXCEPT PT 12, PL S439; CHAPPLE

89.	56042-0190 (LT)

SURFACE RIGHTS ONLY; S 1/2 OF N 1/2 LT 4 CON 1 RICHARDSON EXCEPT PL S391 & PT 4 48R1717; CHAPPLE

90.	56042-0078 (LT)

PCL 21213 SEC RAINY RIVER; N 1/2 OF S 1/2 LT 7 CON 4 RICHARDSON; CHAPPLE

91.	56042-0014 (LT)

PCL 11912 SEC RAINY RIVER; S 1/2 LT 7 CON 2 RICHARDSON EXCEPT PL S439 & A55914 BEING THE MINING RIGHTS; CHAPPLE

92.	56042-0030 (LT)

PCL 14238 SEC RAINY RIVER; N 1/2 LT 8 CON 3 RICHARDSON EXCEPT A55914 BEING THE MINING RIGHTS; CHAPPLE

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93.	56042-0083 (LT)

PCL 22496 SEC RAINY RIVER; N 1/2 OF S 1/2 LT 8 CON 2 RICHARDSON EXCEPT A55914 BEING THE MINING RIGHTS; CHAPPLE

94.	56042-0103 (LT)

PCL 4259 SEC RAINY RIVER; N 1/2 LT 8 CON 1 RICHARDSON EXCEPT PT 4 PL S439, PT 3 48R1985 & A55914 BEING THE MINING RIGHTS; CHAPPLE

95.	56042-0108 (LT)

PCL 4947 SEC RAINY RIVER; S 1/2 LT 8 CON 3 RICHARDSON EXCEPT A55914 BEING THE MINING RIGHTS; CHAPPLE

96.	56042-0122 (LT)

PCL 8070 SEC RAINY RIVER; S 1/2 LT 7 CON 3 RICHARDSON EXCEPT THE E 1/2 & A55914 BEING THE MINING RIGHTS; CHAPPLE

97.	56042-0111 (LT)

PCL 5279 SEC RAINY RIVER; E 1/2 OF S 1/2 LT 8 CON 1 RICHARDSON EXCEPT A62973 BEING THE MINING RIGHTS; CHAPPLE

98.	56032-0280 (LT)

MINING RIGHTS ONLY OF PT MINING CLAIM FF-5877 SENN NOT COVERED BY THE WATERS OF OFF LAKE AS IN SP4199; DISTRICT OF RAINY RIVER

99.	56045-0182 (LT)

MINING RIGHTS ONLY; PT LT 1 CON 2 SIFTON AS IN SLT48145 EXCEPT PT 7, PL S446; UNORGANIZED

100.	56045-0184 (LT)

MINING RIGHTS ONLY; PT LT 1 CON 2 SIFTON AS IN SLT48174 EXCEPT PT 6, PL S446; UNORGANIZED

101.	56045-0186 (LT)

MINING RIGHTS ONLY; PT LT 1 CON 2 SIFTON AS IN SLT48600 EXCEPT PT 3, 5 & 8, PL S446; UNORGANIZED

102.	56045-0180 (LT)

MINING RIGHTS ONLY; S 1/2 LT 2 CON 3 SIFTON; UNORGANIZED

103.	56045-0172 (LT)

MINING RIGHTS ONLY; N 1/2 LT 2 CON 3 SIFTON EXCEPT PT 10, PL S447; UNORGANIZED

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104.	56045-0174 (LT)

MINING RIGHTS ONLY; S 1/2 OF N 1/2 LT 1 CON 3 SIFTON EXCEPT SLT41820 & PT 11, PL S446; UNORGANIZED

105.	56045-0176 (LT)

MINING RIGHTS ONLY; N 1/2 LT 2 CON 2 SIFTON; UNORGANIZED

106.	56045-0178 (LT)

MINING RIGHTS ONLY; S 1/2 LT 1 CON 3 SIFTON EXCEPT PT 9 & 10, PL S446; UNORGANIZED

107.	56041-0159 (LT)

PCL 6721 SEC RAINY RIVER; SW 1/4 SEC 35 TAIT; CHAPPLE

108.	56041-0222 (LT)

SURFACE RIGHTS ONLY; PT SEC 35 TAIT BEING THE NE SUBDIVISION; TOWNSHIP OF CHAPPLE

109.	56041-0221 (LT)

MINING RIGHTS ONLY; PT SEC 35 TAIT BEING THE NE SUBDIVISION TOWNSHIP OF CHAPPLE

110.	56041-0223 (LT)

SURFACE RIGHTS ONLY; NW 1/4 SEC 36 TAIT; CHAPPLE

111.	56041-0224 (LT)

MINING RIGHTS ONLY; NW 1/4 SEC 36 TAIT; CHAPPLE

112.	56041-0225 (LT)

SURFACE RIGHTS ONLY; PT SEC 35 TAIT BEING THE NW 1/4 ; CHAPPLE

113.	56041-0226 (LT)

MINING RIGHTS ONLY; PT SEC 35 TAIT BEING THE NW 1/4 ; CHAPPLE

114.	56041-0215 (LT)

SURFACE RIGHTS ONLY; PT SE 1/4 SEC 35 TAIT PART 1, 48R4044 ; CHAPPLE

115.	56041-0164 (LT)

PCL 4153 SEC RAINY RIVER; PT SEC 36 TAIT BEING THE NE SUBDIVISION EXCEPT PL S390 & PT 5, 6, 8 & 9, 48R1197; CHAPPLE

LOGO

Schedule “B”

Leasehold Lands

1.	56042-0140 (LT)

PCL L1078 SEC RAINY RIVER LEASEHOLD; LT 8 CON 3 RICHARDSON; PT LT 7 CON 3 RICHARDSON BEING THEW 1/2 OF THE S 1/2; MRO; CHAPPLE

2.	56042-0141 (LT)

PCL L1078 SEC RAINY RIVER LEASEHOLD; PT LT 8 CON 2 RICHARDSON BEING ALL OF THE N 1/2 OF THE S 1/2; PT LT 7 CON 2 RICHARDSON BEING THEE 1/2 OF THEN 1/2, & ALL OF THE S 1/2; MRO; CHAPPLE

3.	56042-0142 (LT)

PCL L1078 SEC RAINY RIVER LEASEHOLD; PT LT 8 CON 1 RICHARDSON BEING THEN 1/2; MRO; CHAPPLE

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Schedule “C”

Option Lands

1.	56035-0036 (LT)

PCL 12658 SEC RAINY RIVER; E 1/2 OF S 1/2 LT 9 CON 5 POTTS; CHAPPLE

2.	56035-0089 (LT)

PCL 17905 SEC RAINY RIVER; PT W 1/2 OF S 1/2 LT 9 CON 5 POTTS AS IN SLT74194; CHAPPLE

3.	56035-0168 (LT)

PCL 5301 SEC RAINY RIVER; S 1/2 LT 9 CON 5 POTTS EXCEPT SLT40477, SLT74194 & SLT46701; CHAPPLE

4.	56042-0188 (LT)

SURFACE RIGHTS ONLY; E 1/2 OF S 1/2 LT 9 CON 2 RICHARDSON EXCEPT PT 5 48R1985 & PL S-439; CHAPPLE

5.	56042-0186 (LT)

SURFACE RIGHTS ONLY; E 1/2 OF N 1/2 LT 9 CON 1 RICHARDSON EXCEPT PT 6, PL S439, PT 4 48R1985; CHAPPLE

6.	56042-0172 (LT)

SURFACE RIGHTS ONLY; S 1/2 LT 9 CON 3 RICHARDSON; CHAPPLE

7.	56042-0178 (LT)

SURFACE RIGHTS ONLY; W 1/2 OF N 1/2 LT 12 CON 3 RICHARDSON; CHAPPLE

8.	56042-0176 (LT)

SURFACE RIGHTS ONLY; PT LT 3 CON 1 RICHARDSON AS IN SLT76289; CHAPPLE

9.	56042-0174 (LT)

SURFACE RIGHTS ONLY; S 1/2 LT 11 CON 2 RICHARDSON EXCEPT THE W 1/2 & PT 11 PL S-439; CHAPPLE

10.	56042-0170 (LT)

SURFACE RIGHTS ONLY; S 1/2 LT 12 CON 2 RICHARDSON EXCEPT PT 14, PL S446, PT 17, PL S439, SLT50657; CHAPPLE

11.	56042-0168 (LT)

SURFACE RIGHTS ONLY; N 1/2 LT 12 CON 2 RICHARDSON EXCEPT PT 15, PL S446; CHAPPLE

LOGO

12.	56042-0123 (LT)

PCL 8071 SEC RAINY RIVER; N 1/2 LT 2 CON 3 RICHARDSON; CHAPPLE

13.	56042-0124 (LT)

PCL 8235 SEC RAINY RIVER; N 1/2 LT 3 CON 3 RICHARDSON; CHAPPLE

14.	56032-0281 (LT)

SURFACE RIGHTS ONLY OF PT MINING CLAIM FF-5877 SENN NOT COVERED BY THE WATERS OF OFF LAKE AS IN SP4199; DISTRICT OF RAINY RIVER

15.	56032-0240 (LT)

PCL 9145 SEC RAINY RIVER; SUMMER RESORT LOCATION G2958 SENN ON OFF LAKE EXCEPT A12802; DISTRICT OF RAINY RIVER

16.	56045-0181 (LT)

SURFACE RIGHTS ONLY; PT LT 1 CON 2 SIFTON AS IN SLT48145 EXCEPT PT 7, PL S446; UNORGANIZED

17.	56045-0183 (LT)

SURFACE RIGHTS ONLY; PT LT 1 CON 2 SIFTON AS IN SLT48174 EXCEPT PT 6, PL S446; UNORGANIZED

18.	56045-0185 (LT)

SURFACE RIGHTS ONLY; PT LT 1 CON 2 SIFTON AS IN SLT48600 EXCEPT PT 3, 5, & 8, PL S446; UNORGANIZED

19.	56045-0179 (LT)

SURFACE RIGHTS ONLY; S 1/2 LT 2 CON 3 SIFTON; UNORGANIZED

20.	56045-0171 (LT)

SURFACE RIGHTS ONLY; N 1/2 LT 2 CON 3 SIFTON EXCEPT PT 10, PL S447; UNORGANIZED

21.	56045-0173 (LT)

SURFACE RIGHTS ONLY; S 1/2 OF N 1/2 LT 1 CON 3 SIFTON EXCEPT SLT41820 & PT 11, PL S446; UNORGANIZED

22.	56045-0175 (LT)

SURFACE RIGHTS ONLY; N 1/2 LT 2 CON 2 SIFTON; UNORGANIZED

23.	56045-0177 (LT)

SURFACE RIGHTS ONLY; S 1/2 LT 1 CON 3 SIFTON EXCEPT PT 9 & 10, PL 8446; UNORGANIZED

LOGO

24.	56045-0039 (LT)

PCL 14386 SEC RAINY RIVER; N 1/2 LT 3 CON 2 SIFTON; UNORGANIZED

25.	56045-0098 (LT)

PCL 24968SEC RAINY RIVER; S 1/2 LT 3 CON 3 SIFTON SURFACE RIGHTS ONLY; UNORGANIZED

26.	56041-0163 (LT)

PCL 8386 SEC RAINY RIVER; PT SEC 36 TAIT BEING THE SW SUBDIVISION; CHAPPLE

27.	56041-0219 (LT)

SURFACE RIGHTS ONLY; SE 1/4 SEC 35 TAIT EXCEPT PT 1 48R4044; CHAPPLE

28.	56041-0220 (LT)

MINING RIGHTS ONLY; SE 1/4 SEC 35 TAIT; CHAPPLE

29.	56041-0117 (LT)

PCL 14468 SEC RAINY RIVER; NE 1/4 SEC 22 TAIT; CHAPPLE

30.	56041-0138 (LT)

PCL 9519 SEC RAINY RIVER; NE 1/4 SEC 26 TAIT; CHAPPLE

31.	56041-0140 (LT)

PCL 8719 SEC RAINY RIVER; PT SEC 25 TAIT BEING THE SW SUBDIVISION; CHAPPLE

32.	56041-0023 (LT)

PCL 19642 SEC RAINY RIVER; PT SEC 26 TAIT BEING THE NE 1/4 OF THE SE 1/4; CHAPPLE

33.	56041-0158 (LT)

PCL 14464 SEC RAINY RIVER; PT SEC 34 TAIT BEING THE SE SUBDIVISION EXCEPT SLT68553; CHAPPLE

LOGO

Schedule “D”

Unpatented Claims

See Attached.

LOGO


Mining Claim Client Reports				Page 1 of 4

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KENORA Mining Division - 402288 - RAINY RIVER RESOURCES LTD.


Township/Area	Claim Number	Recording Date	Claim Due Date	Status	Percent Option			Work Required		Total Applied		Total Reserve		Claim Bank
FLEMING	3019809	2004-May-17	2015-May-17	A		100	%	$	4,800	$	43,200	$	377,465	$	0
FLEMING	4211671	2006-Jun-26	2015-Jun-26	A		100	%	$	400	$	2,800	$	0	$	0
FLEMING	4244241	2009-Jan-28	2015-Jan-28	A		100	%	$	6,400	$	25,600	$	25,806	$	0
FLEMING	4244243	2009-Jan-28	2015-Jan-28	A		100	%	$	1,200	$	4,800	$	0	$	0
FLEMING	4245258	2009-Jan-28	2015-Jan-28	A		100	%	$	400	$	1,600	$	0	$	0
FLEMING	4245259	2009-Jan-28	2015-Jan-28	A		100	%	$	800	$	3,200	$	0	$	0
FLEMING	4245260	2009-Jan-28	2014-Jan-28	A		100	%	$	3,200	$	9,600	$	0	$	0
MENARY	4208866	2005-Oct-26	2014-Oct-26	A		100	%	$	6,400	$	44,800	$	0	$	0
MENARY	4208867	2005-Oct-26	2014-Oct-26	A		100	%	$	4,800	$	33,600	$	0	$	0
MENARY	4208868	2005-Oct-26	2014-Oct-26	A		100	%	$	6,400	$	44,800	$	0	$	0
MENARY	4208869	2005-Oct-26	2014-Oct-26	A		100	%	$	6,400	$	44,800	$	0	$	0
MENARY	4208870	2005-Oct-26	2014-Oct-26	A		100	%	$	6,400	$	44,800	$	0	$	0
MENARY	4208871	2005-Oct-26	2014-Oct-26	A		100	%	$	6,000	$	42,000	$	0	$	0
MENARY	4208872	2005-Oct-26	2014-Oct-26	A		100	%	$	6,400	$	44,800	$	5,845	$	0
MENARY	4208873	2005-Oct-26	2014-Oct-26	A		100	%	$	6,400	$	44,800	$	0	$	0
MENARY	4208874	2005-Oct-26	2014-Oct-26	A		100	%	$	6,400	$	44,800	$	0	$	0
MENARY	4208875	2005-Oct-26	2014-Oct-26	A		100	%	$	6,400	$	44,800	$	0	$	0
MENARY	4208876	2005-Oct-26	2014-Oct-26	A		100	%	$	5,600	$	39,200	$	0	$	0
MENARY	4244244	2009-Jan-28	2014-Jan-28	A		100	%	$	4,800	$	14,400	$	0	$	0
MENARY	4244245	2009-Jan-28	2014-Jan-28	A		100	%	$	4,800	$	14,400	$	0	$	0
MENARY	4244247	2009-Jan-28	2014-Jan-28	A		100	%	$	6,400	$	19,200	$	0	$	0


Mining Claim Client Reports				Page 2 of 4


MENARY	4244248	2009-Jan-28	2014-Jan-28	A	100	%	$	6,400	$	19,200	$	0	$	0
POTTS	3012554	2007-Mar-13	2015-Mar-13	A	100	%	$	1,200	$	7,200	$	0	$	0
POTTS	4207826	2006-Feb-20	2015-Feb-20	A	100	%	$	1,600	$	11,200	$	0	$	0
POTTS	4211670	2006-Jun-26	2015-Jun-26	A	100	%	$	1,600	$	11,200	$	0	$	0
POTTS	4211672	2006-Jun-26	2014-Jun-26	A	100	%	$	2,000	$	12,000	$	0	$	0
POTTS	4218605	2007-Apr-19	2014-Apr-19	A	100	%	$	1,600	$	8,000	$	562	$	0
POTTS	4224810	2008-May-06	2013-May-06	A	100	%	$	6,400	$	19,200	$	1,525	$	0
POTTS	4224811	2008-May-06	2013-May-06	A	100	%	$	1,600	$	4,800	$	427	$	0
POTTS	4224812	2008-May-06	2013-May-06	A	100	%	$	4,800	$	14,400	$	2,655	$	0
POTTS	4224813	2008-May-15	2013-May-15	A	100	%	$	800	$	2,400	$	334	$	0
POTTS	4244242	2009-Jan-28	2014-Jan-28	A	100	%	$	2,800	$	8,400	$	0	$	0
POTTS	4245251	2009-Jan-28	2014-Jan-28	A	100	%	$	4,800	$	14,400	$	0	$	0
POTTS	4245252	2009-Jan-28	2014-Jan-28	A	100	%	$	1,942	$	10,858	$	0	$	0
POTTS	4245253	2009-Jan-28	2015-Jan-28	A	100	%	$	3,200	$	12,800	$	0	$	0
POTTS	4245254	2009-Jan-28	2014-Jan-28	A	100	%	$	400	$	1,200	$	0	$	0
POTTS	4245255	2009-Jan-28	2014-Jan-28	A	100	%	$	2,400	$	7,200	$	0	$	0
POTTS	4249688	2010-Mar-01	2015-Mar-01	A	100	%	$	1,600	$	4,800	$	0	$	0
RICHARDSON	1105422	1992-Oct-09	2014-Oct-09	A	100	%	$	864	$	32,736	$	1,583	$	0
RICHARDSON	1105423	1992-Oct-09	2014-Oct-09	A	100	%	$	1,600	$	32,000	$	670	$	0
RICHARDSON	1105425	1992-Oct-09	2014-Oct-09	A	100	%	$	3,200	$	64,000	$	298	$	0
RICHARDSON	1105426	1992-Oct-09	2014-Oct-09	A	100	%	$	800	$	16,000	$	0	$	0
RICHARDSON	1105427	1992-Oct-15	2014-Oct-15	A	100	%	$	1,600	$	32,000	$	0	$	0
RICHARDSON	1105428	1992-Oct-15	2013-Oct-15	A	100	%	$	4,000	$	92,000	$	0	$	0
RICHARDSON	1105430	1992-Oct-15	2013-Oct-15	A	100	%	$	4,000	$	92,000	$	0	$	0
RICHARDSON	1161073	1991-Dec-19	2014-Dec-19	A	100	%	$	3,200	$	67,200	$	1,389	$	0
RICHARDSON	1161074	1991-Dec-19	2014-Dec-19	A	100	%	$	1,600	$	33,600	$	0	$	0
RICHARDSON	1161075	1991-Dec-19	2014-Dec-19	A	100	%	$	800	$	16,800	$	0	$	0
RICHARDSON	1161076	1991-Dec-19	2013-Dec-19	A	100	%	$	4,000	$	96,800	$	1,209	$	0
RICHARDSON	1161079	1991-Dec-19	2014-Dec-19	A	100	%	$	3,200	$	67,200	$	1,456	$	0
RICHARDSON	1161080	1991-Dec-19	2014-Dec-19	A	100	%	$	3,200	$	67,200	$	0	$	0
RICHARDSON	1161081	1991-Dec-19	2014-Dec-19	A	100	%	$	3,200	$	67,200	$	0	$	0


Mining Claim Client Reports				Page 3 of 4


RICHARDSON	1161100	1991-Dec-19	2014-Dec-19	A	100	%	$	3,200	$	67,200	$	0	$	0
RICHARDSON	1161592	1994-Mar-01	2013-Mar-01	A	100	%	$	1,600	$	27,200	$	0	$	0
RICHARDSON	1161604	1994-Mar-01	2013-Mar-01	A	100	%	$	800	$	13,600	$	0	$	0
RICHARDSON	1178215	1995-Feb-24	2013-Feb-24	A	100	%	$	6,400	$	102,400	$	439	$	0
RICHARDSON	1210106	1996-May-27	2015-May-27	A	100	%	$	800	$	13,600	$	425	$	0
RICHARDSON	4251442	2010-Jun-02	2014-Jun-02	A	100	%	$	1,600	$	3,200	$	679	$	0
SENN	3008455	2004-Jun-21	2015-Jun-21	A	100	%	$	5,600	$	50,400	$	0	$	0
SENN	3008456	2004-Jun-21	2015-Jun-21	A	100	%	$	1,600	$	14,400	$	0	$	0
SENN	3012529	2006-Feb-13	2014-Feb-13	A	100	%	$	6,400	$	38,400	$	0	$	0
SENN	3012530	2006-Feb-13	2014-Feb-13	A	100	%	$	6,400	$	38,400	$	0	$	0
SENN	3016066	2006-Feb-13	2014-Feb-13	A	100	%	$	6,400	$	38,400	$	0	$	0
SENN	3016067	2006-Feb-13	2014-Feb-13	A	100	%	$	6,400	$	38,400	$	0	$	0
SENN	3016068	2006-Feb-13	2014-Feb-13	A	100	%	$	6,400	$	38,400	$	0	$	0
SENN	3016069	2006-Feb-13	2015-Feb-13	A	100	%	$	6,400	$	44,800	$	0	$	0
SENN	3016070	2006-Feb-13	2015-Feb-13	A	100	%	$	6,400	$	44,800	$	0	$	0
SENN	4244246	2009-Jan-28	2014-Jan-28	A	100	%	$	5,200	$	15,600	$	0	$	0
SENN	4244249	2009-Jan-28	2014-Jan-28	A	100	%	$	6,400	$	19,200	$	0	$	0
SIFTON	1218904	2012-Jan-09	2014-Jan-09	A	100	%	$	400	$	0	$	0	$	0
TAIT	4253992	2011-Jan-11	2013-Jan-11	A	100	%	$	2,000	$	0	$	277	$	0
TAIT	4253993	2011-Jan-11	2013-Jan-11	A	100	%	$	1,600	$	0	$	869	$	0


Mining Claim Client Reports				Page 4 of 4

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Important Notices

Schedule “D1”

Optioned Unpatented Claims (re English Option Agreements)

Mining Claim No. K 4200490

Mining Claim No. K 4200492

Mining Claim No. K 4200493

Mining Claim No. K 4200494

Mining Claim No. K 4214437

All in the Kenora Mining Division

Optioned Unpatented Claims (re Fred Roisin Option Agreement)

Mining Claim No. K 3016858

Mining Claim No. K 3016859

Mining Claim No. K 3016838

All in the Kenora Mining Division

Optioned Unpatented Claims (re Timberridge Option Agreement)

Mining Claim No. K 3016793

In the Kenora Mining Division

LOGO

Schedule “E”

Qualifications

1.	Any reservations, limitations, provisos and conditions expressed in the original grants from the Crown as the same may be varied by statute.

2.	The exceptions and qualifications contained in section 44(1) of the Land Titles Act (Ontario), paragraphs 7, 8, 9, 10, 12 and 14.

3.	No investigation has been made of the original application for filing in respect of, or the location of the boundaries of any of the Patented Lands, the Leasehold Lands, the Unpatented Claims or the Optioned Unpatented Claims.

4.	No examination of the ground was made to determine if any of the Patented Lands, the Leasehold Lands, the Unpatented Claims or Optioned Unpatented Claims were validly staked and assessment work carried out in compliance with the Mining Act (Ontario) and the regulations thereunder.

5.	The rights, title and interests of the registered or recorded holders and/or Rainy River in and to the Patented Lands, the Leasehold Lands, the Unpatented Claims and Optioned Unpatented Claims are subject to the following:

compliance with the terms of the Mining Act (Ontario) and the regulations pursuant thereto;

the terms of each lease underlying the Leasehold Lands;

the reservations and exceptions contained in the Mining Act (Ontario) and the regulations pursuant thereto, including those noted on the Records;

statutory priorities and preferences and liens, encumbrances or other charges which are extent and are still within the time for recording, or which are valid without recording, in the Mining Recorders Office or Mining Lands Section of the Ontario Ministry of Northern Development and Mines;

any unregistered or unrecorded rights;

we have assumed, without verification, that the grant or creation of any and all rights, title or interests in and to the Patented Lands was in compliance with the Planning Act (Ontario);

the existence of any possible conflict with aboriginal title or rights; and

applicable bankruptcy, insolvency or similar laws affecting creditors’ rights generally.

The enforceability of the Option Agreements may be limited by applicable bankruptcy, insolvency, reorganization, moratorium and other laws affecting creditors’ rights generally and general principles of equity and the discretion that a court of competent jurisdiction may exercise in the granting of equitable remedies.

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NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

APPENDIX B

Rainy River Resources Patented, Leasehold and Mining Claims

(as of March 13, 2013)

Patented/Leased claims on the Rainy River Gold Project property.


	Township/Area	Parcel	PIN (Parcel Identification Number)	Owner and/or Option Name
1	MATHER	17371	56036-0084	RR
2	MATHER	16847	56036-0077	RR
3	POTTS	12658	56035-0036	HUTIKKA
4	POTTS	17905	56035-0089	PETKAU
5	POTTS	5301	56035-0168	PETKAU
6	POTTS	16077	56035-0066	RR
7	POTTS	18689	56035-0098	RR
8	POTTS	5899	56035-0176	RR
9	POTTS	17941	56035-0090	RR
10	POTTS	SRO (former 16784)	56035-0242	RR
11	RICHARDSON	L1078	56042-0140	RR
12	RICHARDSON	L1078	56042-0141	RR
13	RICHARDSON	L1078	56042-0142	RR
14	RICHARDSON	22495 (SRO)	56042-0082	RR
15	RICHARDSON	SRO (former 18580)	56042-0188	BURKELAND/HANN
16	RICHARDSON	MRO (former 18580)	56042-0189	RR
17	RICHARDSON	SRO (former 16342)	56042-0186	BURKELAND
18	RICHARDSON	MRO (former 16342)	56042-0187	RR
19	RICHARDSON	SRO (former 7064)	56042-0160	RR
20	RICHARDSON	MRO (former 7064)	56042-0161	RR
21	RICHARDSON	SRO (former 11087)	56042-0162	RR
22	RICHARDSON	MRO (former 11087)	56042-0163	RR
23	RICHARDSON	25891 (MRO)	56042-0097	RR
24	RICHARDSON	14408 (SRO)	56042-0034	RR
25	RICHARDSON	SRO (former 13275)	56042-0172	ELIUK
26	RICHARDSON	MRO (former 13275)	56042-0173	RR
27	RICHARDSON	26007 (MRO)	56042-0139	RR
28	RICHARDSON	4534 (SRO)	56042-0104	RR
29	RICHARDSON	26007 (MRO)	56042-0102	RR
30	RICHARDSON	5483 (SRO)	56042-0113	RR
31	RICHARDSON	SRO (former 20561)	56042-0178	GERULA, B
32	RICHARDSON	MRO (former 20561)	56042-0179	RR
33	RICHARDSON	SRO (former 18204)	56042-0176	GIBB
34	RICHARDSON	MRO (former 18204)	56042-0177	RR
35	RICHARDSON	25984 (MRO)	56042-0100	RR
36	RICHARDSON	8825 (MRO)	56042-0128	RR
37	RICHARDSON	25892 (MRO)	56042-0098	RR
38	RICHARDSON	SRO (former 4801)	56042-0174	JSD LOGGING
39	RICHARDSON	MRO (former 4801)	56042-0175	RR
40	RICHARDSON	SRO (former 7320)	56042-0170	LEBLANC
41	RICHARDSON	MRO (former 7320)	56042-0171	RR
42	RICHARDSON	SRO (former 7180)	56042-0168	LEBLANC
43	RICHARDSON	MRO (former 7180)	56042-0169	RR
44	RICHARDSON	14407 (SRO)	56042-0033	RR
45	RICHARDSON	25894 (MRO)	56042-0099	RR
46	RICHARDSON	16297	56042-0058	RR
47	RICHARDSON	SRO (former 14196)	56042-0166	RR
48	RICHARDSON	MRO (former 14196)	56042-0167	RR
49	RICHARDSON	MRO (former 10029)	56042-0158	RR
50	RICHARDSON	SRO (former 10029)	56042-0159	RR
51	RICHARDSON	5939	56042-0116	RR


	Township/Area	Parcel	PIN (Parcel Identification Number)	Owner and/or Option Name
52	RICHARDSON	5614	56042-0114	RR
53	RICHARDSON	17725	56042-0063	RR
54	RICHARDSON	17726	56042-0064	RR
55	RICHARDSON	17110	56042-0060	RR
56	RICHARDSON	14462	56042-0036	RR
57	RICHARDSON	MRO (former 10843)	56042-0146	RR
58	RICHARDSON	SRO (former 10843)	56042-0147	RR
59	RICHARDSON	21129	56042-0077	RR
60	RICHARDSON	16779	56042-0145	RR
61	RICHARDSON	16630	56042-0053	RR
62	RICHARDSON	SRO (former 14986)	56042-0164	RR
63	RICHARDSON	MRO (former 14986)	56042-0165	RR
64	RICHARDSON	SRO (former 10961)	56042-0180	RR
65	RICHARDSON	MRO (former 10961)	56042-0181	RR
66	RICHARDSON	SRO (former 9771)	56042-0184	RR
67	RICHARDSON	MRO (former 9771)	56042-0185	RR
68	RICHARDSON	SRO (former 4768)	56042-0182	RR
69	RICHARDSON	MRO (former 4768)	56042-0183	RR
70	RICHARDSON	MRO (former 4950)	56042-0149	RR
71	RICHARDSON	SRO (former 4950)	56042-0148	RR
72	RICHARDSON	15881	56042-0047	RR
73	RICHARDSON	11853	56042-0012	RR
74	RICHARDSON	17392	56042-0062	RR
75	RICHARDSON	14604	56042-0037	RR
76	RICHARDSON	9080	56042-0129	RR
77	RICHARDSON	SRO (former 17154)	56042-0061	RR
78	RICHARDSON	8071	56042-0123	1530600 ONTARIO
79	RICHARDSON	8235	56042-0124	1530600 ONTARIO
80	RICHARDSON	15282	56042-0044	RR
81	RICHARDSON	23322	56042-0088	RR
82	RICHARDSON	17752	56042-0065	RR
83	RICHARDSON	25991 (SRO)	56042-0101	RR
84	RICHARDSON	13137	56042-0021	RR
85	RICHARDSON	10273	56042-0003	RR
86	RICHARDSON	13467	56042-0024	RR
87	RICHARDSON	16307	56042-0050	RR
88	RICHARDSON	16343	56042-0052	RR
89	RICHARDSON	11409 (SRO)	56042-0011	RR
90	RICHARDSON	12324	56042-0018	RR
91	RICHARDSON	22190	56042-0081	RR
92	RICHARDSON	14665	56042-0038	RR
93	RICHARDSON	16820	56042-0056	RR
94	RICHARDSON	10152	56042-0002	RR
95	RICHARDSON	16754	56042-0055	RR
96	RICHARDSON	16956	56042-0059	RR
97	RICHARDSON	14200	56042-0029	RR
98	RICHARDSON	10746	56042-0005	RR
99	RICHARDSON	6520	56042-0117	RR
100	RICHARDSON	5455	56042-0112	RR
101	RICHARDSON	12083	56042-0016	RR
102	RICHARDSON	7654	56042-0121	RR
103	RICHARDSON	9665	56042-0133	RR
104	RICHARDSON	13514	56042-0025	RR


	Township/Area	Parcel	PIN (Parcel Identification Number)	Owner and/or Option Name
105	RICHARDSON	SRO (former 15916)	56042-0190	RR
106	RICHARDSON	21213	56042-0078	RR
107	RICHARDSON	11912 (SRO)	56042-0014	RR
108	RICHARDSON	14238 (SRO)	56042-0030	RR
109	RICHARDSON	4947 (SRO)	56042-0108	RR
110	RICHARDSON	8070 (SRO)	56042-0122	RR
111	RICHARDSON	22496 (SRO)	56042-0083	RR
112	RICHARDSON	4259 (SRO)	56042-0103	RR
113	RICHARDSON	5279 (SRO)	56042-0111	RR
114	SENN	MRO (former 14979)	56032-0280	RR
115	SENN	SRO (former 14979)	56032-0281	KATRIN/STRAND
116	SENN	9145	56032-0240	KATRIN/STRAND
117	SIFTON	SRO (former 13001)	56045-0181	GERULA, M & N
118	SIFTON	MRO (former 13001)	56045-0182	RR
119	SIFTON	SRO (former 13015)	56045-0183	GERULA, M & N
120	SIFTON	MRO (former 13015)	56045-0184	RR
121	SIFTON	SRO (former 13117)	56045-0185	GERULA, M & N
122	SIFTON	MRO (former 13117)	56045-0186	RR
123	SIFTON	SRO (former 8683)	56045-0179	GERULA, B
124	SIFTON	MRO (former 8683)	56045-0180	RR
125	SIFTON	SRO (former 10271)	56045-0171	LEBLANC
126	SIFTON	MRO (former 10271)	56045-0172	RR
127	SIFTON	SRO (former 10798)	56045-0173	LEBLANC
128	SIFTON	MRO (former 10798)	56045-0174	RR
129	SIFTON	SRO (former 13448)	56045-0175	LEBLANC
130	SIFTON	MRO (former 13448)	56045-0176	RR
131	SIFTON	SRO (former 8201)	56045-0177	LEBLANC
132	SIFTON	MRO (former 8201)	56045-0178	RR
133	SIFTON	14386	56045-0039	GERULA, B
134	SIFTON	24968 (SRO)	56045-0098	TIMBERRIDGE
135	TAIT	6721	56041-0159	RR
136	TAIT	SRO (former 5490)	56041-0222	RR
137	TAIT	MRO (former 5490)	56041-0221	RR
138	TAIT	SRO (former 16623)	56041-0223	RR
139	TAIT	MRO (former 16623)	56041-0224	RR
140	TAIT	SRO (former 21172)	56041-0225	RR
141	TAIT	MRO (former 21172)	56041-0226	RR
142	TAIT	8386	56041-0163	TEEPLE, D & J
143	TAIT	SRO – SE ¹/₄ Sec 35 except Pt 1, 48R4044	56041-0219	TEEPLE, W
144	TAIT	MRO – SE ¹/₄ Sec 35	56041-0220	TEEPLE, D & V

145	TAIT	SRO – Pt SE ¹/₄ Sec 35 except Pt 1, 48R4044	56041-0215	RR
146	TAIT	14468	56041-0117	SCHRAM, K
147	TAIT	9519	56041-0138	BRAGG, D
148	TAIT	8719	56041-0140	BRAGG, D & J
149	TAIT	19642	56041-0023	BRAGG/PAQUETTE
150	TAIT	7153	56041-0164	RR
151	TAIT	14464	56041-0158	ROISIN

Unpatented mining claims on the Rainy River Gold Project property. All claims are active and owned 100 percent by Rainy River Resources Ltd.


	Township/Area	Claim Number	Recording Date		Claim Due Date			Work Required		Total Applied		Total Reserve
1	FLEMING	3019809		2004-May-17		2015-May-17		$	4,800	$	43,200	$	377,465
2	FLEMING	4211671		2006-Jun-26		2015-Jun-26		$	400	$	2,800	$	0
3	FLEMING	4244241		2009-Jan-28		2015-Jan-28		$	6,400	$	25,600	$	25,806
4	FLEMING	4244243		2009-Jan-28		2015-Jan-28		$	1,200	$	4,800	$	0
5	FLEMING	4245258		2009-Jan-28		2015-Jan-28		$	400	$	1,600	$	0
6	FLEMING	4245259		2009-Jan-28		2015-Jan-28		$	800	$	3,200	$	0
7	FLEMING	4245260		2009-Jan-28		2014-Jan-28		$	3,200	$	9,600	$	0
8	MENARY	4208866		2005-Oct-26		2014-Oct-26		$	6,400	$	44,800	$	0
9	MENARY	4208867		2005-Oct-26		2014-Oct-26		$	4,800	$	33,600	$	0
10	MENARY	4208868		2005-Oct-26		2014-Oct-26		$	6,400	$	44,800	$	0
11	MENARY	4208869		2005-Oct-26		2014-Oct-26		$	6,400	$	44,800	$	0
12	MENARY	4208870		2005-Oct-26		2014-Oct-26		$	6,400	$	44,800	$	0
13	MENARY	4208871		2005-Oct-26		2014-Oct-26		$	6,000	$	42,000	$	0
14	MENARY	4208872		2005-Oct-26		2014-Oct-26		$	6,400	$	44,800	$	5,845
15	MENARY	4208873		2005-Oct-26		2014-Oct-26		$	6,400	$	44,800	$	0
16	MENARY	4208874		2005-Oct-26		2014-Oct-26		$	6,400	$	44,800	$	0
17	MENARY	4208875		2005-Oct-26		2014-Oct-26		$	6,400	$	44,800	$	0
18	MENARY	4208876		2005-Oct-26		2014-Oct-26		$	5,600	$	39,200	$	0
19	MENARY	4244244		2009-Jan-28		2014-Jan-28		$	4,800	$	14,400	$	0
20	MENARY	4244245		2009-Jan-28		2014-Jan-28		$	4,800	$	14,400	$	0
21	MENARY	4244247		2009-Jan-28		2014-Jan-28		$	6,400	$	19,200	$	0
22	MENARY	4244248		2009-Jan-28		2014-Jan-28		$	6,400	$	19,200	$	0
23	POTTS	3012554		2007-Mar-13		2015-Mar-13		$	1,200	$	7,200	$	0
24	POTTS	4207826		2006-Feb-20		2015-Feb-20		$	1,600	$	11,200	$	0
25	POTTS	4211670		2006-Jun-26		2015-Jun-26		$	1,600	$	11,200	$	0
26	POTTS	4211672		2006-Jun-26		2014-Jun-26		$	2,000	$	12,000	$	0
27	POTTS	4218605		2007-Apr-19		2014-Apr-19		$	1,600	$	8,000	$	562
28	POTTS	4224810		2008-May-06		2013-May-06	**	$	6,400	$	19,200	$	1,525
29	POTTS	4224811		2008-May-06		2013-May-06	**	$	1,600	$	4,800	$	427
30	POTTS	4224812		2008-May-06		2013-May-06	**	$	4,800	$	14,400	$	2,655
31	POTTS	4224813		2008-May-15		2013-May-15	**	$	800	$	2,400	$	334
32	POTTS	4244242		2009-Jan-28		2014-Jan-28		$	2,800	$	8,400	$	0
33	POTTS	4245251		2009-Jan-28		2014-Jan-28		$	4,800	$	14,400	$	0
34	POTTS	4245252		2009-Jan-28		2014-Jan-28		$	1,942	$	10,858	$	0
35	POTTS	4245253		2009-Jan-28		2015-Jan-28		$	3,200	$	12,800	$	0
36	POTTS	4245254		2009-Jan-28		2014-Jan-28		$	400	$	1,200	$	0
37	POTTS	4245255		2009-Jan-28		2014-Jan-28		$	2,400	$	7,200	$	0
38	POTTS	4249688		2010-Mar-01		2015-Mar-01		$	1,600	$	4,800	$	0
39	RICHARDSON	1105422		1992-Oct-09		2014-Oct-09		$	864	$	32,736	$	1,583
40	RICHARDSON	1105423		1992-Oct-09		2014-Oct-09		$	1,600	$	32,000	$	670
41	RICHARDSON	1105425		1992-Oct-09		2014-Oct-09		$	3,200	$	64,000	$	298
42	RICHARDSON	1105426		1992-Oct-09		2014-Oct-09		$	800	$	16,000	$	0
43	RICHARDSON	1105427		1992-Oct-15		2014-Oct-15		$	1,600	$	32,000	$	0
44	RICHARDSON	1105428		1992-Oct-15		2013-Oct-15	**	$	4,000	$	92,000	$	0
45	RICHARDSON	1105430		1992-Oct-15		2013-Oct-15	**	$	4,000	$	92,000	$	0
46	RICHARDSON	1161073		1991-Dec-19		2014-Dec-19		$	3,200	$	67,200	$	1,389
47	RICHARDSON	1161074		1991-Dec-19		2014-Dec-19		$	1,600	$	33,600	$	0
48	RICHARDSON	1161075		1991-Dec-19		2014-Dec-19		$	800	$	16,800	$	0
49	RICHARDSON	1161076		1991-Dec-19		2013-Dec-19	**	$	4,000	$	96,800	$	1,209
50	RICHARDSON	1161079		1991-Dec-19		2014-Dec-19		$	3,200	$	67,200	$	1,456


	Township/Area	Claim Number	Recording Date		Claim Due Date			Work Required		Total Applied		Total Reserve
51	RICHARDSON	1161080		1991-Dec-19		2014-Dec-19		$	3,200	$	67,200	$	0
52	RICHARDSON	1161081		1991-Dec-19		2014-Dec-19		$	3,200	$	67,200	$	0
53	RICHARDSON	1161100		1991-Dec-19		2014-Dec-19		$	3,200	$	67,200	$	0
54	RICHARDSON	1161592		1994-Mar-01		2013-Mar-01	**	$	1,600	$	27,200	$	0
55	RICHARDSON	1161604		1994-Mar-01		2013-Mar-01	**	$	800	$	13,600	$	0
56	RICHARDSON	1178215		1995-Feb-24		2013-Feb-24	**	$	6,400	$	102,400	$	439
57	RICHARDSON	1210106		1996-May-27		2015-May-27		$	800	$	13,600	$	425
58	RICHARDSON	4251442		2010-Jun-02		2014-Jun-02		$	1,600	$	3,200	$	679
59	SENN	3008455		2004-Jun-21		2015-Jun-21		$	5,600	$	50,400	$	0
60	SENN	3008456		2004-Jun-21		2015-Jun-21		$	1,600	$	14,400	$	0
61	SENN	3012529		2006-Feb-13		2014-Feb-13		$	6,400	$	38,400	$	0
62	SENN	3012530		2006-Feb-13		2014-Feb-13		$	6,400	$	38,400	$	0
63	SENN	3016066		2006-Feb-13		2014-Feb-13		$	6,400	$	38,400	$	0
64	SENN	3016067		2006-Feb-13		2014-Feb-13		$	6,400	$	38,400	$	0
65	SENN	3016068		2006-Feb-13		2014-Feb-13		$	6,400	$	38,400	$	0
66	SENN	3016069		2006-Feb-13		2015-Feb-13		$	6,400	$	44,800	$	0
67	SENN	3016070		2006-Feb-13		2015-Feb-13		$	6,400	$	44,800	$	0
68	SENN	4244246		2009-Jan-28		2014-Jan-28		$	5,200	$	15,600	$	0
69	SENN	4244249		2009-Jan-28		2014-Jan-28		$	6,400	$	19,200	$	0
70	SIFTON	1218904		2012-Jan-09		2014-Jan-09		$	400	$	0	$	0
71	TAIT	4253992		2011-Jan-11		2013-Jan-11	**	$	2,000	$	0	$	277
72	TAIT	4253993		2011-Jan-11		2013-Jan-11	**	$	1,600	$	0	$	869


**Note:		As at March 11, 2013. MNDM website indicates that claim is “ACTIVE – WORK REPORT PENDING”.

Unpatented mining claims of the “English Option” agreement.


	Township/Area	Claim Number	Recording Date	Claim Due Date	Work Required		Total Applied		Total Reserve
1	TAIT	4200492	2006-Oct-27	2014-Oct-27	$	1,600	$	9,600	$	78,536
2	TAIT	4200494	2006-Oct-27	2014-Oct-27	$	5,200	$	31,200	$	94,336
3	TAIT	4214437	2010-Mar-03	2015-Mar-03	$	1,600	$	4,800	$	41,783
4	TAIT	4200490	2006-Oct-27	2014-Oct-27	$	800	$	4,800	$	0
5	TAIT	4200493	2006-Oct-27	2014-Oct-27	$	3,600	$	21,600	$	86,914

Unpatented mining claims of the “Roisin Option” agreement.


	Township/Area	Claim Number	Recording Date	Claim Due Date	Work Required		Total Applied		Total Reserve
1	POTTS	3016858	2010-Jul-08	2016-Jul-08	$	1,600	$	6,400	$	0
2	RICHARDSON	3016838	2010-Jul-08	2013-Jul-08	$	3,200	$	3,200	$	0
3	RICHARDSON	3016859	2010-Jul-08	2014-Jul-08	$	1,600	$	3,200	$	0

Unpatented mining claims of the “Timberridge Option” agreement.


	Township/Area	Claim Number	Recording Date	Claim Due Date	Work Required		Total Applied		Total Reserve
1	SIFTON	3016793	2009-Sep-25	2013-Sept-25	$	1,600	$	3,200	$	1,165

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

APPENDIX C

Nuinsco Exploration Activities (1993 to 2002)

Summary of Nuinsco Exploration Activities at the Rainy River Gold Project (1993 – 2002)


Activity ¹	Date	Performed by:
Rotasonic drilling	26/06/93 - 01/07/93	Midwest Drilling
IP survey	December, 1993	Val d’Or Géophysique
Magnetometer survey	December, 1993	Val d’Or Géophysique
Landsat linear study	August, 1993	DOZ Consulting Group
Reconnaissance mapping and sampling	04/07/93 - 27/08/93	Nuinsco Resources
Rotasonic drilling	March, 1994	Midwest Drilling
Reverse circulation drilling	06/03/94 - 01/04/94	Bradley Bros. - Overburden Drilling
Diamond drilling	08/11/94 - 20/12/94	Ultra Mobile Diamond Drilling
Grid mapping and sampling	May-June, 1994	Nuinsco Resources
Soil Sampling/Enzyme Leach	June-August, 1994	Nuinsco Resources
Reverse circulation drilling	Winter, 1995	Bradley Bros. - Overburden Drilling
Diamond drilling	04/01/95 - 16/12/95	Ultra Mobile Diamond Drilling
IP survey	25/10/95 - 10/12/95	JVX Geophysics
Trenching and stripping, mapping	Field season, 1995	Nuinsco Resources
Soil Sampling/Enzyme Leach	Field season, 1995	Nuinsco Resources
Reverse circulation drilling	Winter, 1996	Bradley Bros. - Overburden Drilling
Diamond drilling	Throughout year, 1996	Ultra Mobile Diamond Drilling
Diamond drilling	26/01/96 - 22/07/96	Bradley Brothers Diamond Drilling
UTEM survey	28/03/96 - 23/04/96	Lamontagne Geophysics
Surface pulse EM	June, 1996	Crone Geophysics
Surface pulse EM	25/09/96 - 18/12/96	JVX Geophysics
Borehole pulse EM	17/06/96 - 10/08/96	Crone Geophysics
Borehole pulse EM	26/09/96 - 21/10/96	Crone Geophysics
Borehole pulse EM	25/09/96 - 18/12/96	JVX Geophysics
IP survey	25/09/96 - 18/12/96	JVX Geophysics
Magnetometer survey	25/09/96 - 18/12/96	JVX Geophysics
Outcrop stripping	October, 1996	Nuinsco Resources
Reverse circulation drilling	Winter, 1997	Bradley Bros. - Overburden Drilling
Reverse circulation drilling	September, October, 1997	Bradley Bros. - Overburden Drilling
Diamond drilling	Throughout year, 1997	Ultra Mobile Diamond Drilling
Diamond drilling	22/01/97 - 07/04/97	Bradley Brothers Diamond Drilling
Airborne EM and Magnetic survey	04/10/97 - 08/10/97	Geoterrex-Dighem
Surface pulse EM	December, 1997	Crone Geophysics
Borehole pulse EM	03/17/97 - 03/19/97	Crone Geophysics
Borehole pulse EM	September, December 1997	Crone Geophysics
IP survey	28/09/97 - 06/10/97	Quantec IP
Local detailed mapping	May-September, 1997	Nuinsco Resources
Outcrop stripping	October, 1997	Nuinsco Resources
Surface PEM survey	Jan. Feb., 1998	Crone Geophysics
Diamond drilling	01/04/98 - 15/03/98	Ultra Mobile Diamond Drilling.
Reverse circulation drilling	Jan - Feb 1998	Bradley Bros.- Overburden Drilling
Line cutting/Magnetometer survey	March, 1998	Mtec Geophysics Inc.
Diamond Drilling	04/01/98 - 28/04/98	Ultra Mobile Diamond Drilling
Diamond Drilling	15/06/99 - 19/07/99	Ultra Mobile Diamond Drilling
Diamond Drilling	08/08/99 - 01/09/99	Bradley Brothers Diamond Drilling
Airborne EM and Magnetic Survey	15-19 August 2000	Aeroquest Limited
Geochemical Compilation	2000-2001	Franklin Geoscience and Nuinsco Personnel
Magnetotelluric Geophysical Survey	2001-2002	Phoenix Geophysics
Mapping/Prospecting	May 2001 - Sept 2001	Nuinsco Resources
Diamond Drilling	15/11/01 - 18/02/02	Diamond Drilling, Bradley Brothers

¹	From Mackie et al. 2003

NI 43-101 Technical Report

Feasibility Study of the Rainy River Gold Project

APPENDIX D

Analytical Quality Control Data and Relative Precision Charts

(December 2011 – July 2012)

Time Series Plots for Field Blank and Certified Standard Samples Assayed by ALS Chemex Laboratories during December 2011 to July 2012 – Gold Assays

LOGO

Time Series Plots for Field Blank and Certified Standard Samples Assayed by ALS Chemex Laboratories during December 2011 to July 2012 – Gold Assays

LOGO

Time Series Plots for Field Blank and Certified Standard Samples Assayed by ALS Chemex Laboratories during December 2011 to July 2012 – Silver Assays

LOGO

Bias Charts, Quantile-Quantile and Relative Precision Plots for Field Duplicate Samples Assayed by ALS Chemex Laboratories during December 2011 and July 2012 – Gold Assays

LOGO

Bias Charts, Quantile-Quantile and Relative Precision Plots for Check Assay Samples (ALS Chemex Laboratories versus Activation Laboratories) – Gold Assays

LOGO