EX-99.1

Prepared by:

In Collaboration with:

EX-99.1 2 d675939dex991.htm EX-99.1 EX-99.1

Exhibit 99.1

NI 43-101

Feasibility Study of the Rainy River Project,

Ontario, Canada

LOGO

Prepared for:

LOGO

Suite 1800, Two Bentall Centre

555 Burrard Street

Vancouver, BC

Report Date: February 14, 2014

Effective Date: January 16, 2014

3098007-AD0000-30-AET-0001-00

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.

Colm Keogh, P. Eng., AMC Mining Consultants (Canada), Ltd.

Mo Molavi, P. Eng., AMC Mining Consultants (Canada), Ltd.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Project

DATE AND SIGNATURE PAGE

This Report is effective as of the 16^th day of January 2014.


Original Signed				February 14, 2014
Colin Hardie, P.Eng.				Date
Department Manager, Mining and Metals
BBA Inc.

Original Signed				February 14, 2014
David Runnels, Eng.				Date
Project Manager, Mining and Metals
BBA Inc.

Original Signed				February 14, 2014
Patrice Live, Eng.				Date
Manager of Mining
BBA Inc.

Original Signed				February 14, 2014
Sheila E. Daniel, M.Sc., P.Geo.				Date
Head Environmental Management
Senior Associate Geoscientist
AMEC – Environment & Infrastructure

Original Signed				February 14, 2014
David G. Ritchie, P.Eng.				Date
Geotechnical Engineering Group Manager
Senior Associate Geotechnical Engineer
AMEC – Environment & Infrastructure

Original Signed				February 14, 2014
Adam Coulson, PhD., P.Eng.				Date
Rock Engineering Group Manager
Senior Associate Rock Mechanics Engineer
AMEC – Environment & Infrastructure

Original Signed				February 14, 2014
Glen Cole, P.Geo.				Date
Principal Consultant (Resource Geology)
SRK Consulting (Canada) Inc.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Project


Original Signed				February 14, 2014
Dorota El-Rassi, P.Eng.				Date
Senior Consultant (Resource Geology)
SRK Consulting (Canada) Inc.

Original Signed				February 14, 2014
Colm Keogh, P.Eng.				Date
Principal Mining Engineer
AMC Mining Consultants (Canada) Ltd.

Original Signed				February 14, 2014
Mo Molavi, P.Eng.				Date
Principal Mining Engineer
Mining Services Manager
AMC Mining Consultants (Canada) Ltd.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Project

LOGO

CERTIFICATE OF QUALIFIED PERSON

This certificate applies to the technical report prepared for New Gold Inc. (“New Gold”) entitled: “Feasibility Study of the Rainy River Project, Ontario, Canada” signed on February 14, 2014 (the “Technical Report”) and effective January 16, 2014.

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

I am currently employed as a Department Manager and Metallurgist 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;

I visited the Rainy River Project 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;

I am independent of the issuer as independence is described in Section 1.5 of NI 43-101;

I am responsible for Sections 1, 2, 3, 4, 19, 21 (except 21.4.4, 21.4.5, 21.5.2, 21.5.5, 21.6.4, and 21.6.5), 22, 23, 25, 26, 27 and Appendices A and 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; the “Preliminary Economic Assessment Update of the Rainy River Gold Property” prepared by BBA in October 2012 and the “Feasibility Study of the Rainy River Gold Project” prepared by BBA in May 2013 and re- addressed to New Gold in July 2013;

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

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LOGO

CERTIFICATE OF QUALIFIED PERSON

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

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;

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;

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;

I visited the Rainy River Project 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;

I am independent of the issuer as independence is described in Section 1.5 of NI 43-101;

I am responsible for Sections 13, 17, 18 (except 18.1.3, 18.8, 18.9 and 18.10), 24 and Appendix F of the Technical Report;

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)

NI 43-101 Technical Report

Feasibility Study of the Rainy River Project

LOGO

CERTIFICATE OF QUALIFIED PERSON

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

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;

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;

I visited the Rainy River Project 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;

I am independent of the issuer as independence is described in Section 1.5 of NI 43-101;

I am responsible for Sections 15 (except 15.3), 16 (except 16.2.1, 16.3, 16.4 and 16.5), 21.4.4, 21.4.5, 21.5.2 and 21.6.4 and contributed to Sections 1, 25 and 26 of the 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

10)

NI 43-101 Technical Report

Feasibility Study of the Rainy River Project

LOGO

CERTIFICATE OF QUALIFIED PERSON

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

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;

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;

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

I visited the Rainy River Project site on May 19, 2011;

I am independent of the issuer as independence is described in Section 1.5 of NI 43-101;

I am responsible for Section 20 and 18.10 and contributed to Sections 1, 25 and 26 of the Technical Report;

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)

NI 43-101 Technical Report

Feasibility Study of the Rainy River Project

LOGO

CERTIFICATE OF QUALIFIED PERSON

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

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

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.

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

I visited the Rainy River Project site on September 4, 2013;

I am independent of the issuer as independence is described in Section 1.5 of NI 43-101;

I am responsible for Sections 16.2.1.2, 18.1.3, 18.8 and 18.9 and contributed to Sections 1, 25 and 26 of the Technical Report;

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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LOGO

CERTIFICATE OF QUALIFIED PERSON

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

I am currently employed as a Senior Associate Rock Mechanics Engineer in the consulting firm AMEC Earth & Infrastructure, a division of AMEC Americas Ltd.:

160 Traders Blvd. E.,

Suite 110

Mississauga, Ontario L4Z 3K7 Canada

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;

I visited the Rainy River Project site on September 25 to 27, 2013;

I am independent of the issuer as independence is described in Section 1.5 of NI 43-101;

I am responsible for Sections 16.2.1.1 and 16.3.6 and contributed to Sections 1, 25 and 26 of the Technical Report;

I have had prior involvement with the subject property having co-authored a previous technical report entitled “Feasibility Study of the Rainy River Gold Project” prepared by BBA in May 2013 and re-addressed to New Gold in July 2013;

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

10)

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LOGO

CERTIFICATE OF QUALIFIED PERSON

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

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;

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);

I have personally inspected the Rainy River Project site and surrounding areas on a few occasions, most recently from April 30 to May 2, 2013;

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;

I am independent of the issuer as independence is described in Section 1.5 of NI 43-101;

I am responsible for sections 5, 6, 7, 8, 9, 10, 11, 12 and Appendices C to E and contributed to Sections 1, 25 and 26 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; the “Preliminary Economic Assessment Update of the Rainy River Gold Property” prepared by BBA in October 2012 and the “Feasibility Study of the Rainy River Gold Project” prepared by BBA in May 2013 and re-addressed to New Gold in July 2013;

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 Project

LOGO

CERTIFICATE OF QUALIFIED PERSON

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

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;

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;

I have not personally visited the Rainy River Project site 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;

I am independent of the issuer as independence is described in Section 1.5 of NI 43-101;

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 and the “Feasibility Study of the Rainy River Gold Project” prepared by BBA in May 2013 and re-addressed to New Gold in July 2013;

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 Project

LOGO

CERTIFICATE OF QUALIFIED PERSON

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

I am currently employed as a Principal Mining Engineer in the consulting firm AMC Mining Consultants (Canada) Inc. with a business address at Suite 202, 200 Granville Street, Vancouver, British Columbia, V6C 1S4.

I graduated with a Bachelor of Applied Science degree (Mining) from the University of British Columbia in Vancouver, British Columbia in 1988;

I am a chartered member of the Association of Professional Engineers and Geoscientists of British Columbia. I have practiced my profession continuously since 1988 in a range of operational, technical and consulting roles;

I visited the Rainy River Project site on September 26, 2013;

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;

I am independent of the issuer as independence is described in Section 1.5 of NI 43-101;

I am responsible for Sections 15.3, 16.3 (except 16.3.6), 21.5.5 and 21.6.5 and contributed to sections 1.15.2, 1.15.3, 1.16.4, 1.16.5, 25.5, 26 and 26.2 of the Technical Report;

I have no prior involvement with the Property that is the subject of the 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

10)

NI 43-101 Technical Report

Feasibility Study of the Rainy River Project

LOGO

CERTIFICATE OF QUALIFIED PERSON

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

I am a Principal Mining Engineer with AMC Mining Consultants (Canada) Ltd with a business address at Suite 202, 200 Granville Street, Vancouver, British Columbia, V6C 1S4.

I am a graduate Laurentian University (B. Eng. in Mining Engineering, 1979) and McGill University (M. Eng. in Mining Engineering specializing in Rock Mechanics and Mining Methods, 1987).

I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia, License #37594 and a member of the Canadian Institute of Mining, Metallurgy and Petroleum. I have worked as a Mining Engineer for a total of 34 years since my graduation from university and have relevant experience in project management, feasibility studies and technical report preparations for mining projects in North America. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”).

I did not complete a personal inspection of the Rainy River Project site but relied on a site visit completed by Mr. Colm Keogh, P.Eng., a co-author of the Technical Report;

I am independent of the issuer as independence is described in Section 1.5 of NI 43-101;

I am responsible for Sections 16.4 and 16.5 and contributed to section 1.16.3 of the Technical Report;

I have no prior involvement with the Property that is the subject of the 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

10)

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CAUTIONARY NOTE WITH RESPECT TO FORWARD LOOKING INFORMATION

Certain information and statements contained in this report are “forward looking” in nature. All information and statements in this report, other than statements of historical fact, that address events, results, outcomes or developments that New Gold and/or the Qualified Persons who authored this report expect to occur are “forward-looking statements”. Forward-looking statements are statements that are not historical facts and are generally, but not always, identified by the use of forward-looking terminology such as “plans”, “expects”, is “expected”, “budget”, “scheduled”, “estimates”, “forecasts”, “intends”, “anticipates”, “projects”, “potential”, “believes” or variations 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” or the negative connotation of such terms. Forward-looking statements include, but are not limited to, statements with respect to the economic and feasibility parameters of the Rainy River project: the cost and timing of the development of the project; the proposed mine plan and mining method, stripping ratio, processing method and rates and production rates; grades; projected metallurgical recovery rates; infrastructure, capital, operating and sustaining costs; the projected life of mine and other expected attributes of the Rainy River project; the NPV and IRR and payback period of capital; cash costs and all-in sustaining costs; the success and continuation of exploration activities; estimates of mineral reserves and resources; the future price of gold; the timing of the environmental assessment process; 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.

All forward-looking statements in this report are necessarily based on opinions and estimates made as of the date such statements are made and are subject to important risk factors and uncertainties, many of which cannot be controlled or predicted. Material assumptions regarding forward-looking statements are discussed in this report, where applicable. In addition to, and subject to, such specific assumptions discussed in more detail elsewhere in this report, the forward-looking statements in this report are subject to the following assumptions: (1) there being no signification disruptions affecting the development and operation of the project; (2) the exchange rate between the Canadian dollar and U.S. dollar being approximately consistent with current levels; (3) the availability of certain consumables and services and the prices for diesel, natural gas, fuel oil, electricity and other key supplies being approximately consistent with current levels; (4) labour and materials costs increasing on a basis consistent with current expectations; (5) permitting and arrangements with First Nations and other Aboriginal groups being consistent with current expectations; (6) that all environmental approvals, required permits, licenses and authorizations will be obtained from the relevant

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governments and other relevant stakeholders within the expected timelines; (7) certain tax rates, including the allocation of certain tax attributes to the project; (8) the availability of financing for New Gold’s development activities; (9) the timelines for exploration and development activities on the project; and (10) assumptions made in mineral resource and reserve estimates, including geological interpretation grade, recovery rates, gold price assumption, and operational costs; and general business and economic conditions.

Forward-looking statements involve known and unknown risks, uncertainties and other factors which may cause the actual results, performance or achievements to be materially different from any of the 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 and economic parameters discussed herein not being realized; decrease of future gold prices; cost of labour, supplies, fuel and equipment rising; actual results of current exploration; adverse changes in project parameters including discrepancies between actual and estimated production, reserves, resources and recoveries; exchange rate fluctuations; delays and costs inherent in consulting and accommodating rights of First Nations and other Aboriginal groups; title risks; regulatory risks, and political or economic developments in Canada; changes to tax rates; changes to New Afton’s mine plan or profitability or to New Gold’s asset profile that might alter the allocation of tax attributes to Rainy River; risks and uncertainties with respect to obtaining necessary permits and necessary surface rights, land use rights and other tenure from the Crown and private landowners or delays in obtaining same; risks associated with maintaining and renewing permits and complying with permitting requirements, including without limitation approval of the environmental assessment and receipt of all related permits, authorizations or other rights, including under the Endangered Species Act and Public Lands Act; and other risks involved in the gold exploration and development industry; as well as those risk factors discussed elsewhere in this report, in New Gold’s latest Annual Information Form, Management’s Discussion and Analysis and its other SEDAR filings from time to time. All forward-looking statements herein are qualified by this cautionary statement. Accordingly, readers should not place undue reliance on forward-looking statements. New Gold and the Qualified Persons who authored 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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CAUTIONARY NOTE TO U.S. READERS CONCERNING ESTIMATES OF MINERAL RESERVES

AND MINERAL RESOURCES

Information concerning the Rainy River Project has been prepared in accordance with Canadian standards under applicable Canadian securities laws, and may not be comparable to similar information for United States companies. The terms “Mineral Resource”, “Measured Mineral Resource”, “Indicated Mineral Resource” and “Inferred Mineral Resource” used in this Report are Canadian mining terms as defined in the Canadian Institute of Mining, Metallurgy and Petroleum (“CIM”) Definition Standards for Mineral Resources and Mineral Reserves adopted by CIM Council on November 27, 2010 and incorporated by reference in National Instrument 43-101 (“NI 43-101”). While the terms “Mineral Resource”, “Measured Mineral Resource”, “Indicated Mineral Resource” and “Inferred Mineral Resource” are recognized and required by Canadian securities regulations, they are not defined terms under standards of the United States Securities and Exchange Commission. As such, certain information contained in this Report concerning descriptions of mineralization and resources under Canadian standards is not comparable to similar information made public by United States companies subject to the reporting and disclosure requirements of the United States Securities and Exchange Commission.

An “Inferred Mineral Resource” has a greater amount of uncertainty as to its existence and as to its economic and legal feasibility. It cannot be assumed that all or any part of an “Inferred Mineral Resource” will ever be upgraded to a higher confidence category.Readers are cautioned not to assume that all or any part of an “Inferred Mineral Resource” exists or is economically or legally mineable.

Under United States standards, mineralization may not be classified as a “Reserve” unless the determination has been made that the mineralization could be economically and legally produced or extracted at the time the Reserve estimation is made. Readers are cautioned not to assume that all or any part of the Measured or Indicated Mineral Resources that are not Mineral Reserves will ever be converted into Mineral Reserves. In addition, the definitions of “Proven Mineral Reserves” and “Probable Mineral Reserves” under CIM standards differ in certain respects from the standards of the United States Securities and Exchange Commission.

NI 43-101 Technical Report

Feasibility Study of the Rainy River Project

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		1-2

1.4	Property Description		1-4

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

1.6	Project History		1-5

1.7	Geological Setting and Mineralization		1-7

1.8	Deposit Types		1-7

1.9	Exploration		1-8

1.10	Drilling		1-8

1.11	Sampling Method, Approach and Analyses		1-9

1.12	Data Verification		1-9

1.13	Metallurgical Test work		1-10

	1.13.1	Historical Test work	1-10

	1.13.2	Mineralogy	1-10

	1.13.3	Flowsheet Determination Test work	1-11

	1.13.4	Comminution Tests	1-12

	1.13.5	Gravity Separation	1-13

	1.13.6	Cyanide Leaching	1-13

	1.13.7	Cyanide Destruction Test work	1-14

	1.13.8	Environmental Test work	1-14

	1.13.9	Test work 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		1-24

	1.16.1	Open Pit Operations	1-25

	1.16.2	Open Pit Production Schedule	1-26

	1.16.3	Underground Mine Infrastructure	1-27

	1.16.4	Underground Operations	1-28

	1.16.5	Underground Development and Production Schedule	1-30

	1.16.6	Proposed Overall Mine Plan (Open Pit and Underground)	1-31

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1.17	Process Plant		1-33

1.18	Project Infrastructure		1-36

	1.18.1	Site Water Management	1-39

	1.18.2	Tailings Management Area	1-40

1.19	Market Studies and Contracts		1-40

1.20	Environmental and Permitting		1-40

1.21	Capital and Operating Costs		1-42

	1.21.1	Capital Costs	1-42

	1.21.2	Operating Costs	1-45

1.22	Economic Analysis		1-51

1.23	Adjacent Properties		1-56

1.24	Other Relevant Data and Information		1-57

1.25	Conclusions		1-57

1.26	Recommendations and Future Work Program		1-58

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		2-4

2.5	Site Visit		2-5

2.6	Acknowledgement		2-5

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	Environmental Approvals 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

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6.	HISTORY		6-1

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	7-6

	7.2.2	Structural Geology	7-8

7.3	Mineralization		7-15

	7.3.1	Auriferous Sulphide and Quartz-Sulphide Stringers and Veins in Felsic Quartz-Phyric Rocks	7-16

	7.3.2	Deformed Quartz-Ankerite-Pyrite Shear Veins in Mafic Volcanic Rocks	7-19

	7.3.3	Silver-Rich Deformed Sulphide-Quartz Veins within Tuffaceous Rocks	7-20

	7.3.4	Nickel-Copper-PGE Mineralization	7-21

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 (2005-2013)		9-1

10.	DRILLING		10-1

10.1	Drilling from 2004 to 2013		10-1

	10.1.1	Introduction	10-1

	10.1.2	Drilling Procedures	10-3

10.2	Drilling Pattern and Density		10-5

10.3	SRK Comments		10-5

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

11.1	Sampling Method and Approach		11-1

11.2	Sample Preparation and Analyses		11-2

	11.2.1	Nuinsco Samples	11-2

	11.2.2	Rainy River Samples	11-2

	11.2.3	Metallurgical Testing	11-6

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11.3	Specific Gravity Data		11-7

11.4	Quality Assurance and Quality Control Programs		11-10

11.5	SRK Comments		11-11

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	12-5

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	Variability Sample Selection	13-3

	13.2.2	Sample Characterization	13-5

13.3	Mineralogy		13-7

13.4	Flowsheet Selection Testwork		13-8

	13.4.1	Flotation Option	13-9

	13.4.2	Gravity and Gravity Tailings Whole Rock Leach	13-16

	13.4.3	Flowsheet Selection	13-18

13.5	Comminution Tests		13-18

	13.5.1	Crusher Work Index	13-18

	13.5.2	Unconfined Compressive Strength	13-19

	13.5.3	JK Drop Weight and SAG Mill Comminution	13-19

	13.5.4	SAGDesign	13-22

	13.5.5	Bond Work Index	13-24

	13.5.6	ModBond and A x b	13-27

	13.5.7	Bond Abrasion Index	13-28

	13.5.8	Grinding Circuit Design	13-29

13.6	Gravity Separation		13-31

	13.6.1	Gravity Recoverable Gold	13-31

	13.6.2	Gravity Variability Testwork	13-32

13.7	Heap Leaching		13-34

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13.8	Cyanide Leaching		13-35

	13.8.1	Gravity Tailings Leaching	13-35

	13.8.2	Cyanide Concentration	13-38

	13.8.3	Pre-Conditioning	13-39

	13.8.4	Oxygen (O₂) vs. Air	13-40

	13.8.5	Lead Nitrate Addition	13-40

	13.8.6	Intrepid Zone Leaching Kinetics	13-41

	13.8.7	Grade Recovery Variability Tests	13-43

	13.8.8	Diagnostic Leach Testwork	13-46

	13.8.9	Mercury Assays	13-47

13.9	Cyanide Destruction Testwork		13-47

13.10	Carbon-in-Pulp Modelling		13-48

13.11	Thickener Sizing Testwork		13-50

	13.11.1	Flocculant Screening	13-50

	13.11.2	Sedimentation Testwork	13-51

13.12	Rheology		13-52

13.13	Linear Screen Sizing Testwork		13-53

13.14	Environmental Testwork		13-54

13.15	Gold and Silver Recovery Curves		13-55

13.16	Testwork Interpretation		13-57

14.	MINERAL RESOURCE ESTIMATION		14-1

14.1	Introduction		14-1

14.2	Resource Estimation Procedures		14-2

14.1	Resource Database		14-2

14.3	Solid Body Modelling		14-4

	14.3.1	Introduction	14-4

	14.1.1	The ODM/17 Zone	14-7

	14.1.2	The 433, HS and New Zones	14-8

14.4	Compositing		14-10

14.5	Evaluation of Outliers		14-11

14.6	Statistical Analysis and Variography		14-14

	14.6.1	Statistical Analysis	14-14

	14.6.2	Variography	14-24

14.7	Block Model and Grade Estimation		14-30

	14.7.1	Block Model Definition	14-30

	14.7.2	Grade and Specific Gravity Estimation	14-30

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14.8	Model Validation and Sensitivity		14-38

14.9	Mineral Resource Classification		14-39

14.10	Mineral Resource Statement		14-40

14.11	Grade Sensitivity Analysis		14-48

14.12	Previous Mineral Resource Estimates		14-51

15.	MINERAL RESERVE ESTIMATE		15-1

15.1	Introduction		15-1

15.2	Open Pit Mining		15-1

	15.2.1	Resource Block Model	15-1

	15.2.2	Open Pit Optimization	15-4

	15.2.3	Detailed Mine Design	15-8

	15.2.4	In-Pit Dilution and Mine Recovery	15-20

	15.2.5	Open Pit Mineral Reserves	15-21

15.3	Underground Mining		15-22

	15.3.1	Underground Mineral Reserves	15-22

	15.3.2	Mining Shapes	15-27

	15.3.3	Dilution and Recovery Estimates	15-27

	15.3.4	Recovery Factors	15-30

	15.3.5	Mineral Reserves	15-30

15.4	Open Pit and Underground Mineral Reserves		15-30

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	Material Management	16-15

	16.2.4	Open Pit Mine Equipment and Operations	16-21

	16.2.5	Open Pit Mine Personnel Requirements	16-33

16.3	Underground Mining Methods		16-37

	16.3.1	Mine Design	16-38

	16.3.2	Stope Design	16-44

	16.3.3	Mining Method and Sequence	16-46

	16.3.4	Waste Management and Stope Filling	16-51

	16.3.5	Development and Production Schedule	16-54

	16.3.6	Geotechnical	16-58

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	16.3.7	Drill and Blast	16-63

	16.3.8	Mobile Equipment Requirements	16-65

	16.3.9	Ventilation	16-70

	16.3.10	Emergency Preparedness	16-78

16.4	Underground infrastructure		16-80

	16.4.1	Refuge Stations	16-80

	16.4.2	Mine Dewatering	16-81

	16.4.3	Compressed Air	16-83

	16.4.4	Mine Water Supply	16-83

	16.4.5	Fuelling and Lubrication	16-84

	16.4.6	Workshop Facility	16-85

	16.4.7	Explosives Magazine	16-86

	16.4.8	Underground (CAF Loading Station)	16-87

	16.4.9	Communications and Automation	16-88

	16.4.10	Electrical Distribution	16-90

16.5	Underground Manpower Requirements		16-92

	16.5.1	Manpower	16-92

	16.5.2	Schedule	16-92

	16.5.3	Organization	16-92

16.6	Combined Production Schedule		16-94

17.	RECOVERY METHODS		17-1

17.1	Proposed Process Flowsheet		17-1

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

	17.2.1	Primary Crushing	17-6

	17.2.2	Crushed Rock Handling and Storage	17-7

	17.2.3	Processing Plant and Tailings Handling	17-7

	17.2.4	Primary and Secondary Grinding	17-9

	17.2.5	Gravity Circuit	17-10

	17.2.6	Cyanide Leaching Circuit	17-10

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

	17.2.8	Tailings Management and Cyanide Destruction	17-12

	17.2.9	Refining Area and Gold Room	17-13

	17.2.10	Reagent Areas	17-13

	17.2.11	Control Room and Maintenance Shop	17-13

	17.2.12	Offices and Change House	17-14

	17.2.13	Metallurgical Laboratory	17-14

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17.3	Energy, Water and Consumable Requirements		17-14

	17.3.1	Energy	17-14

	17.3.2	Water	17-14

	17.3.3	Consumables	17-16

18.	PROJECT INFRASTRUCTURE		18-1

18.1	General Site Works		18-1

	18.1.1	Primary Site and Access Roads	18-1

	18.1.2	Mine Haul Roads	18-2

	18.1.3	Geotechnical	18-2

18.2	Mine Services Facilities		18-3

18.3	General Offices and Assay Laboratory		18-4

	18.3.1	Main Administration Building	18-5

	18.3.2	Mine Office and Dry	18-5

	18.3.3	Plant Office	18-5

18.4	Parking Area		18-6

18.5	Assay Lab		18-6

18.6	Fuel Storage and Dispensing		18-6

18.7	Electrical and Communication		18-7

	18.7.1	Emergency Power	18-8

	18.7.2	Communication	18-9

18.8	Tailings Management Area		18-10

	18.8.1	Tailings Deposition Plan	18-10

	18.8.2	Dam Design	18-11

	18.8.3	Construction and Operational Considerations	18-13

	18.8.4	Site Water Management	18-16

	18.8.5	Water Management Structures	18-17

	18.8.6	Runoff and Seepage Collection	18-18

	18.8.7	Process Plant Water Supply - Preparations for Start-up	18-21

	18.8.8	Process Plant Water Supply - Operations	18-22

18.9	Water Treatment Design Basis and Operation		18-23

	18.9.1	Water Treatment Design and Operation	18-24

	18.9.2	Water Treatment Plant Opportunities	18-26

18.10	Fish Compensation Works		18-27

	18.10.1 Background		18-27

	18.10.2 Proposed Offset Measures		18-29

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19.	MARKET STUDIES AND CONTRACTS		19-1

19.1	Market Studies		19-1

19.2	Commodity Price Projections		19-1

19.3	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

	20.2.3	Comments on the Project	20-7

20.3	Environmental Studies		20-10

	20.3.1	Overview	20-10

	20.3.2	Climate, Air Quality and Sound	20-12

	20.3.3	Physiography, Soils and Geology	20-13

	20.3.4	Hydrology and Hydrogeology	20-15

	20.3.5	Surface Water, Sediment and Groundwater Quality	20-17

	20.3.6	Biological Environment - Existing Conditions	20-19

	20.3.7	Human Environment	20-22

20.4	Traditional Knowledge (“TK”) and Traditional Land Use (“TLU”)		20-24

20.5	Cultural Heritage Resources		20-25

20.6	Environmental Sensitivities		20-26

20.7	Regulatory Context		20-27

	20.7.1	Current Regulatory Status	20-27

	20.7.2	Environmental Approvals Required for Proposed Operations	20-27

20.8	Preliminary Environmental Impact		20-32

20.9	Effects Analysis		20-33

	20.9.1	Air Quality and Sound	20-33

	20.9.2	Streamflow, Aquatic Habitats and Species	20-34

	20.9.3	Groundwater	20-36

	20.9.4	Vegetation and Terrestrial Habitat	20-36

	20.9.5	Terrestrial and Avian Species	20-37

	20.9.6	Species at Risk	20-37

	20.9.7	Traditional Land Use	20-38

	20.9.8	Socio-economic	20-38

	20.9.9	Human Health	20-38

	20.9.10 Cultural Heritage Resources		20-38

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20.10	Reclamation Approach		20-39

	20.10.1	Open Pit and Underground Mine	20-39

	20.10.2	Stockpiles	20-40

	20.10.3	Tailings Management Area	20-41

	20.10.4	Aggregate Sources	20-41

	20.10.5	Buildings, Machinery, Equipment and Infrastructure	20-42

	20.10.6	Petroleum Products, Chemicals and Explosives	20-42

	20.10.7	Roads, Pipelines and Power Distribution	20-42

	20.10.8	Site Drainage and Water Structures	20-43

	20.10.9	Waste Management	20-43

	20.10.10	Offsite Facilities	20-43

21.	CAPITAL AND OPERATING COSTS		21-1

21.1	Capital Costs - Introduction		21-1

	21.1.1	Assumptions	21-1

	21.1.2	Exclusions	21-2

21.2	Capital Cost Summary		21-2

21.3	Basis of Estimate		21-4

21.4	Pre-Production Capital Costs		21-11

	21.4.1	Introduction	21-11

	21.4.2	Overhead Power Line	21-11

	21.4.3	Highway 600 Re-alignment	21-11

	21.4.4	Open Pit Overburden Pre-Stripping, Waste Removal and Ore Stockpiling	21-12

	21.4.5	Open Pit Mining Equipment	21-12

	21.4.6	Site Development and Process Facilities	21-14

	21.4.7	Tailings, Water Management and Treatment	21-15

	21.4.8	Indirect Costs	21-16

	21.4.9	Owner’s Costs	21-19

	21.4.10	Contingency	21-19

21.5	Sustaining Capital Costs		21-20

	21.5.1	Introduction	21-20

	21.5.2	Open Pit Mine Equipment	21-20

	21.5.3	Mobile Equipment	21-21

	21.5.4	Site Development	21-22

	21.5.5	Underground Mine	21-22

	21.5.6	Tailings Water Management and Treatment	21-27

	21.5.7	Fish Habitat Compensation Costs	21-28

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	21.5.8	Chapple Township Compensation Costs	21-28

	21.5.9	Salvage Value	21-28

	21.5.10 Reclamation and Closure Costs		21-28

21.6	Operating Costs		21-29

	21.6.1	Introduction	21-29

	21.6.2	Power and Fuel	21-32

	21.6.3	Total Employees	21-32

	21.6.4	Open Pit Operating Costs	21-33

	21.6.5	Underground Operating Costs	21-36

	21.6.6	Process Plant	21-38

	21.6.7	G&A Costs	21-42

	21.6.8	Royalties	21-43

	21.6.9	Transportation and Refining	21-43

22.	ECONOMIC ANALYSIS		22-1

22.1	Introduction		22-1

22.2	Methods, Assumptions and Basis		22-1

22.3	Royalties		22-4

22.4	Salvage Value		22-4

22.5	Taxation		22-4

22.6	Financial Analysis Summary		22-6

22.7	Sensitivity Analysis		22-8

23.	ADJACENT PROPERTIES		23-1

23.1	Introduction		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

24.3	Hazardous Waste Management		24-1

24.4	Execution Strategy		24-2

24.5	Management Procedures		24-2

24.6	EPCM Project Controls and Reporting		24-4

24.7	Project Scheduling		24-5

24.8	Procurement and Contracts		24-7

24.9	Site Development		24-7

24.10	Construction		24-7

	24.10.1 Construction Management Responsibilities		24-7

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	24.10.2	Construction Power	24-8

	24.10.3	Construction Labour Requirement	24-8

24.11	Process Facilities		24-9

	24.11.1	Critical Path and Installation Methodology	24-9

24.12	Tailings Management Area (‘‘TMA’’) Earthworks		24-10

24.13	Commissioning		24-10

24.14	Mechanical Completion		24-11

24.15	Risk Management		24-11

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, Processing and General & Administrative		25-7

25.7	Infrastructure		25-9

25.8	Environmental Permitting		25-10

25.9	Financial Analysis		25-10

25.10	Conclusion		25-11

26.	RECOMMENDATIONS		26-1

26.1	Proposed 2014 Work Program Budget		26-1

26.2	2014 Detailed Recommendations		26-2

27.	REFERENCES		27-1


APPENDIX A –		Rainy River Resources Title Opinion (November 27, 2013)

APPENDIX B –		Rainy River Resources Patented, Leasehold and Mining Claims (November 22, 2013)

APPENDIX C –		Nuinsco Exploration Activities (1993 to 2002)

APPENDIX D –		Analytical Quality Control Data and Relative Precision Charts (December 2011 - July 2013)

APPENDIX E –		Domain Variograms

APPENDIX F –		Rainy River Project Site Plan

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LIST OF FIGURES


Figure 1-1:	Phase I, Phase II and Final Pit	1-27

Figure 1-2:	Isometric View of the Rainy River Underground Mine	1-28

Figure 1-3:	Typical Stope Long Section and Cross Sections	1-29

Figure 1-4:	Combined Gold and Silver Production	1-32

Figure 1-5:	Schematic Process Flowsheet	1-35

Figure 1-6:	General Site Layout	1-37

Figure 1-7:	Annual Operating Cash Costs (USD/oz. Au) with Silver Credit	1-50

Figure 1-8:	Life-of-Mine Cash Flow Projections	1-54

Figure 1-9:	Pre-Tax Net Present Value (NPV) Sensitivity Analysis at 5% Discount Rate	1-55

Figure 4-1:	Location of Rainy River Project (as of November 22, 2013)	4-3

Figure 4-2:	Land Tenure Map of the Rainy River Gold Project (as of November 22, 2013)	4-5

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

Figure 7-1:	Regional Bedrock Geology of the Area West of Fort Frances (modified from Percival and Easton, 2007)	7-3

Figure 7-2:	Bedrock Geological Interpretation for the Area Surrounding the Rainy River Project (from Rainy River, 2013)	7-5

Figure 7-3:	Regional Structural Trends on the Rainy River Gold Project, Interpreted from Aeromagnetics (Rankin, 2013)	7-9

Figure 7-4:	Structural Fabrics Affecting Rock Types	7-11

Figure 7-5:	Evidence for Strike-Slip Kinematics of Late Brittle Faulting (SRK, 2011)	7-12

Figure 7-6:	Pressure Shadows Around Rigid Objects in Dacitic Rock	7-13

Figure 7-7:	Sulphide Mineralization Deformed by Folding in Core from the Rainy River Project (SRK, 2011)	7-14

Figure 7-8:	Structural Control of the Plunge of the Gold Mineralization at the Rainy River Project	7-15

Figure 7-9:	ODM/17 Zone Gold Mineralization (SRK, 2011)	7-17

Figure 7-10:	ODM/17 High-Grade Gold Mineralization (SRK, 2011)	7-17

Figure 7-11:	433 High-Grade Gold Mineralization (SRK, 2011)	7-18

Figure 7-12:	Higher Grade Gold Mineralization - Borehole NR10-474 from 188.0 to 234.0 m. (SRK, 2011)	7-20

Figure 7-13:	Intrepid Zone Gold Mineralization. Deformed Pyrite-Sphalerite Veins and Stringers Borehole NR131542 (Rainy River, 2013)	7-21

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)	8-2

Figure 8-2:	Schematic Geological Setting and Hydrothermal Alteration Associated with Gold-rich Volcanogenic Hydrothermal Systems (after Hannington et al., 1999)	8-4

Figure 8-3:	Schematic Section and Hydrothermal Alteration Associated with the Rainy River Project (Sparkes & Wartman, 2012)	8-5

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Figure 9-1:	Section Showing the Footwall Silver Zone below the Previous PEA Starter Pit (December 23, 2011) (from Rainy River	9-3

Figure 9-2:	Section through the Intrepid Zone	9-4

Figure 10-1:	Core Drilling Data by Period (1994 to August 16^th, 2013)	10-2

Figure 10-2:	Drill Collar Plan in Relation to Resource Domains within the Main Rainy River Area and Showing Conceptual Pit Outline	10-2

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)	11-8

Figure 13-1:	Sample Locations for Comminution Variability Testwork	13-3

Figure 13-2:	Sample Locations for Metallurgical Variability Testwork	13-4

Figure 13-3:	Sample Locations for Comminution (Left) and Leaching (Right) Variability Testwork (Cross- Section View, Looking West)	13-4

Figure 13-4:	Specific Energy vs. Particle Size for IsaMill Testing	13-14

Figure 13-5:	Specific Energy vs. Particle Size for SMD Testing	13-16

Figure 13-6:	SMC Data Distribution with JK DWT Calibration Points	13-21

Figure 13-7:	Full Bond Ball Mill Work Indices vs. ModBond Work Indices (200 Mesh)	13-25

Figure 13-8:	Distribution of ModBond Indices (200 Mesh) by Zone	13-27

Figure 13-9:	ModBond Wi vs. A x b	13-28

Figure 13-10:	Gold Gravity Recovery vs. Head Grade	13-33

Figure 13-11:	Silver Gravity Recovery vs. Head Grade	13-33

Figure 13-12:	Heap Leach Gold Recovery Curve	13-34

Figure 13-13:	Gravity Tailings Leach Residue vs. Grind Size	13-36

Figure 13-14:	Cost and Revenue Analysis by Grind Size	13-37

Figure 13-15:	Gold Recovery vs. Time at Different NaCN Concentrations	13-38

Figure 13-16:	Gold and Silver Cyanide Leaching Kinetics (Intrepid Zone)	13-42

Figure 13-17:	Gold Residue vs. Head Grade (Variability Tests)	13-45

Figure 13-18:	Silver Residue vs. Head Grade (Variability Tests)	13-46

Figure 13-19:	CIP Modeling Isotherms	13-49

Figure 13-20:	Initial Pit Composite Yield Stress vs. Solids Density	13-52

Figure 13-21:	Remaining-Life-of-Mine Composite Yield Stress vs. Solids Density	13-53

Figure 14-1:	Location of New Boreholes Drilled on the Main Rainy River Deposit	14-5

Figure 14-2:	Location of Drill Collars and High Grade Subdomain for the Intrepid Zone	14-6

Figure 14-3:	Isometric View of the Rainy River Mineralization Wireframes Modelled by SRK with Borehole Data (View looking towards the West)	14-6

Figure 14-4:	Histogram Distribution of Raw Sample Lengths	14-11

Figure 14-5:	Cumulative Frequency Plot for Gold Composites	14-13

Figure 14-6:	Examples of Variogram Models for Rainy River Deposit	14-25

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Figure 14-7:	Cross-Section 425,775E Comparing Blocks Populated with Gold Grades and Informing Data	14-38

Figure 14-8:	Cross-Section 425, 335E Comparing Blocks Populated With Gold Grades and Informing Data in the Intrepid Zone	14-39

Figure 14-9:	Schematic Vertical Section Illustrating Criteria Considered for Preparing the Mineral Resource Statement for the Rainy River Project (View Looking East)	14-43

Figure 14-10:	Rainy River Project Global Grade Tonnage Curves (Open Pit and Underground Material Combined)	14-48

Figure 14-11:	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:	Rainy River Theoretical Pit Shell and Bayfield constraint (Plan View)	15-7

Figure 15-3:	Detailed Open Pit Mine Design (Plan View)	15-10

Figure 15-4:	Final Pit Design and LG Optimization - Isometric View	15-11

Figure 15-5:	Final Pit Design and LG Optimization - Elevation 290 masl	15-12

Figure 15-6:	Final Pit Design and LG Optimization - Elevation 160 masl	15-13

Figure 15-7:	Final Pit Design and LG Optimization - Elevation 10 masl	15-14

Figure 15-8:	Final Pit Design and LG Optimization - Cross Section (East 425 400, looking West)	15-15

Figure 15-9:	Final Pit Design and LG Optimization - Cross Section (East 425 500, looking West)	15-16

Figure 15-10:	Final Pit Design and LG Optimization - Cross Section (East 425 600, looking West)	15-17

Figure 15-11:	Final Pit Design and LG Optimization - Cross Section (East 425 700, looking West)	15-18

Figure 15-12:	Final Pit Design and LG Optimization - Cross Section (East 425 800, looking West)	15-19

Figure 15-13:	Isometric View of the Rainy River Underground Mine	15-23

Figure 15-14:	Typical ODM Main Zone Cross-Section	15-23

Figure 15-15:	Typical Intrepid Zone Cross-Sections	15-24

Figure 15-16:	Unplanned Dilution from Hangingwall and Footwall Rock	15-28

Figure 15-17:	Backfill Dilution	15-29

Figure 16-1:	Isometric View of the Rainy River Open Pit and Underground Mine, Looking North	16-1

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

Figure 16-3-	Phase 1, Phase 2 and Final Pit - 3D View	16-7

Figure 16-4:	Phase 1, Phase 2 and Final Pit - Cross-section (East 425 500, Looking West)	16-7

Figure 16-5:	Phase 1, Phase 2 and Final Pit - Cross-section (East 425 700, Looking West)	16-8

Figure 16-6:	Open Pit Mine Planning - End of 2015 (Year -2)	16-9

Figure 16-7:	Open Pit Mine Planning - End of 2016 (Year -1)	16-10

Figure 16-8:	Open Pit Mine Planning - End of 2017 (Year 1)	16-11

Figure 16-9:	Open Pit Mine Planning - End of 2018 (Year 2)	16-11

Figure 16-10:	Open Pit Mine Planning - End of 2020 (Year 4)	16-12

Figure 16-11:	Open Pit Mine Planning - End of 2022 (Year 6)	16-12

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Figure 16-12:	Final Pit Contour - End of 2025 (Year 9)	16-13

Figure 16-13:	Open Pit Material Movement over the Life-of-Mine	16-14

Figure 16-14:	Site Plan Showing West and East Mine Rock Stockpiles	16-16

Figure 16-15:	East Mine Rock Stockpile Area	16-16

Figure 16-16:	West Mine Rock Stockpile Area	16-17

Figure 16-17:	In-Pit Dumping Area and Dumping Sequence	16-20

Figure 16-18:	Cycle Time Trend over LOM	16-28

Figure 16-19:	LOM Haul Truck Fleet	16-29

Figure 16-20:	Sublevel Arrangement in Cross-section	16-40

Figure 16-21:	Typical Level Plan	16-40

Figure 16-22:	Standard Design - Ore Drive	16-43

Figure 16-23:	Standard Design - Main Decline, Internal Ramps and Level Accesses	16-43

Figure 16-24:	Mineable Stope Shape - Intrepid Zone	16-45

Figure 16-25:	Mineable Stope Shape - ODM and 17 East Zones	16-45

Figure 16-26:	Mining Zones	16-46

Figure 16-27:	Typical Stope Long Section	16-48

Figure 16-28:	Typical Stope Cross-sections	16-48

Figure 16-29:	CAF System Schematic	16-51

Figure 16-30:	Extent of Mine Development at the Onset of Production (April 2019)	16-55

Figure 16-31:	Underground Production Profile	16-57

Figure 16-32:	Production by Activity	16-57

Figure 16-33:	View North West of the Underground Mine Geometry Developed for the Map3D Numerical Stress Modelling, Indicating Zones and Main Underground Boreholes (AMEC, 2013G)	16-58

Figure 16-34:	Modified Stability Graph for ODM West Zone above 500 m depth Indicating Stability of Stope Surfaces Based on Potential Design Limits and Actual Final Stope Dimensions	16-60

Figure 16-35:	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)	16-62

Figure 16-36:	Longitudinal Downhole Stope Long Section	16-64

Figure 16-37:	Drop Raise Drillhole Pattern for the Longitudinal Downhole Stope (Section View)	16-64

Figure 16-38:	Rainy River Ventilation System	16-72

Figure 16-39:	General Arrangement Refuge Station Plan	16-81

Figure 16-40:	Typical Sump Flow Diagram	16-82

Figure 16-41:	Dewatering Level Sump and Pumping	16-82

Figure 16-42:	General Arrangement Underground Main Sump Layout	16-83

Figure 16-43:	Typical Pressure Reducing Valve (‘‘PRV”) Station	16-84

Figure 16-44:	Underground Fuel Consumption Profile	16-85

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Figure 16-45:	General Arrangement - Underground Fuel Station	16-85

Figure 16-46:	Underground Maintenance Workshops Layout	16-86

Figure 16-47:	ANFO Storage Plan and Sections	16-87

Figure 16-48:	Cap & Powder Magazine Plan and Sections	16-87

Figure 16-49:	Underground Cemented Rock Fill Loading Station	16-88

Figure 16-50:	Underground Power Requirement Profile	16-91

Figure 16-51:	Manpower Loading By Year	16-94

Figure 16-52:	Annual Concentrator Feed	16-95

Figure 16-53:	Annual Concentrator Feed	16-96

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 Pond Water Balance	17-16

Figure 18-1:	Typical Cross Section - TMA South Dam Section (Refer to Table 18-8 for the construction fill material descriptions)	18-14

Figure 18-2:	Typical Cross Section - TMA West and North Sections (Refer to Table 18-8 for the construction fill material descriptions)	18-15

Figure 18-3:	Altered/Displaced Waters Frequented by Fish	18-28

Figure 21-1:	Project Operating Costs	21-29

Figure 21-2:	Annual Operating Cash Costs (USD/oz. Au) with Silver Credit	21-32

Figure 21-3:	Total Personnel	21-33

Figure 22-1:	Life-of-Mine Cash Flow Projection	22-7

Figure 22-2:	Life-of-Mine Cash Flow Projection (Pre-tax and After-tax, discount rate: 5%)	22-8

Figure 22-3:	Sensitivity of the Net Present Value (Pre-tax) to Selected Financial Variables	22-10

Figure 24-1:	Project Management Organization Chart	24-3

Figure 24-2:	Project Milestones	24-6

Figure 24-3:	Monthly Construction Manpower Graph	24-9

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LIST OF TABLES


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

Table 1-2:	Key Outcomes - Gold Production¹	1-3

Table 1-3:	Key Outcomes - Silver Production¹	1-3

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

Table 1-5:	Consolidated Mineral Resource Statement, Rainy River Gold Project, SRK Consulting (Canada) Inc., November 2, 2013^1,2,3,4,5	1-18

Table 1-6:	Pit Optimization Parameters¹	1-20

Table 1-7:	Open Pit and Underground Proven and Probable Mineral Reserves^1,2,3,4,5,6	1-23

Table 1-8:	Open Pit Mine Schedule	1-26

Table 1-9:	Underground Production Schedule	1-31

Table 1-10:	Total Milled from OP and UG	1-33

Table 1-11:	Overall Project Capital Cost Summary	1-43

Table 1-12:	Key Project Operating Costs	1-45

Table 1-13:	Open Pit Operating Costs	1-46

Table 1-14:	LOM Underground Operating Costs by Activity	1-48

Table 1-15:	LOM Underground Mine General Cost Breakdown	1-48

Table 1-16:	Total and All-in Sustaining Cash Costs	1-50

Table 1-17:	Financial Analysis Summary⁵	1-53

Table 1-18:	Selected Sensitivities, Pre-Tax NPV at 5% Discount Rate (CAD $M)	1-55

Table 1-19:	Sensitivity Results for Metal Price and Exchange Rate Variations, $CAD	1-56

Table 1-20:	Sensitivity Results for Metal Price and Exchange Rate Variations, $USD	1-56

Table 1-21:	Rainy River Project Development Activities	1-57

Table 1-22:	Proven and Probable Reserves (November 2, 2013)	1-58

Table 1-23:	Budget for 2014 (US Dollars)	1-59

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 Statement¹ for the Rainy River Project, Ontario, Mackie et al., December 23, 2003.	6-3

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

Table 6-3:	Mineral Resource Statement¹ for the Rainy River Project, Ontario, SRK Consulting (Canada) Inc., April 28, 2009	6-5

Table 6-4:	Mineral Resource Statement¹, Rainy River Project, Ontario, SRK Consulting (Canada) Inc., February 26, 2010	6-6

Table 6-5:	Mineral Resource Statement¹, Rainy River Project, Ontario, SRK Consulting (Canada) Inc., February 24, 2011	6-7

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Table 6-6:	Mineral Resource Statement¹, Rainy River Project, Ontario, SRK Consulting (Canada) Inc., June, 29 2011	6-8

Table 6-7:	Mineral Resource Statement¹, Rainy River Project, Ontario,SRK Consulting (Canada) Inc., February, 24 2012	6-9

Table 6-8:	Consolidated Mineral Resource Statement¹, Rainy River Project, Ontario SRK Consulting (Canada) Inc., October 10, 2012	6-10

Table 9-1:	Summary of Exploration Work by Rainy River on the Rainy River Project between 2005 and 2013	9-6

Table 10-1:	Core Drilling Completed on the Rainy River Project (1994-2013)¹	10-1

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

Table 11-2:	Summary Statistics of Specific Gravity Data for the Intrepid Zone	11-9

Table 12-1:	Summary of Analytical Quality Control Data Produced between December 2011 and July 2012 for the Main Rainy River Deposit	12-3

Table 12-2:	Summary of Analytical Quality Control Data Produced between January 2012 to June 2013 for the Intrepid Zone	12-4

Table 13-1:	Master Composite Proportions	13-1

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

Table 13-3:	Testwork Sample Head Grades	13-6

Table 13-4:	Rougher Flotation Results	13-9

Table 13-5:	Cleaner Flotation Results	13-10

Table 13-6:	Flotation Concentrate Leaching Results (Gold Assays)	13-11

Table 13-7:	Flotation Concentrate Leaching Results (Silver Assays)	13-11

Table 13-8:	Flotation Tailings Leaching Results	13-12

Table 13-9:	IsaMill Testwork Results	13-13

Table 13-10:	Stirred Media Detritor Testwork Results	13-15

Table 13-11:	Gravity Tailings Leach Results (Gold)	13-17

Table 13-12:	Gravity Tailings Leach Results (Silver)	13-17

`Table 13-13:	Crusher Work Index Results	13-18

Table 13-14:	JK Drop Weight and Corresponding SMC Results	13-20

Table 13-15:	SAG Mill Comminution SMC and M_ia Values	13-22

Table 13-16:	SAGDesign Testwork Results	13-23

Table 13-17:	Comparison between SAGDesign and JK DWT Results	13-24

Table 13-18:	Bond and ModBond Results	13-26

Table 13-19:	Abrasion Index Results	13-28

Table 13-20:	SAG and Ball Mill Simulation Results¹	13-30

Table 13-21:	Gravity Recoverable Gold Results	13-32

Table 13-22:	Additional Gravity Tailings Leaching Results (Gold Assays)	13-35

Table 13-23:	Additional Gravity Tailings Leaching Results (Silver Assays)	13-36

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Table 13-24:	Cyanide Concentration Testwork Results	13-38

Table 13-25:	Pre-Conditioning vs. No Pre-Conditioning Testwork Results	13-39

Table 13-26:	O₂ vs. Air and Lead Nitrate Addition Testwork Results	13-40

Table 13-27:	Intrepid Zone Leaching Kinetics Tests (Gold)	13-41

Table 13-28:	Intrepid Zone Leaching Kinetics Tests (Silver)	13-42

Table 13-29:	Gold Leaching Variability Testwork Average Results	13-43

Table 13-30:	Silver Leaching Variability Testwork Average Results	13-44

Table 13-31:	Cyanide Destruction Testwork Results	13-48

Table 13-32:	Flocculant Description	13-50

Table 13-33:	Sedimentation Testwork Results	13-51

Table 13-34:	Linear Screen Sizing Testwork Results	13-53

Table 13-35:	Summary of Geochemical Environmental Testing	13-54

Table 13-36:	Residue and Gravity Recovery Curves	13-56

Table 13-37:	Simplified Gold and Silver Recovery Curves	13-56

Table 13-38:	Gold and Silver Recoveries vs. Head Grade	13-57

Table 13-39:	Average Yearly Gold and Silver Recoveries	13-59

Table 14-1:	Rock Codes in the Rainy River Project Block Model	14-5

Table 14-2:	Summary of Metal Capping Levels Applied to each Resource Domain	14-12

Table 14-3:	Basic Statistics for Gold Composites for All Resource Domains	14-15

Table 14-4:	Basic Statistics for Capped Gold Composites for All Resource Domains	14-16

Table 14-5:	Basic Statistics for Silver Composites for All Resource Domains	14-17

Table 14-6:	Basic Statistics for Capped Silver Composites for All Resource Domains	14-18

Table 14-7:	Basic Statistics of Composites for Domain 200 (Zone 34)	14-19

Table 14-8:	Basic Statistics for Calcium Uncapped Composites for All Resource Domains	14-20

Table 14-9:	Basic Statistics for Calcium Capped Composites for All Resource Domains	14-21

Table 14-10:	Basic Statistics for Sulphur Uncapped Composites for All Resource Domains	14-22

Table 14-11:	Basic Statistics for Sulphur Capped Composites for All Resource Domains	14-23

Table 14-12:	Modeled Gold Variogram Parameters for All Resource Domains Grade Interpolation	14-26

Table 14-13:	Modeled Silver Variogram Parameters for All Resource Domains Grade Interpolation	14-27

Table 14-14:	Modeled Calcium Variogram Parameters for All Resource Domains Grade Interpolation	14-28

Table 14-15:	Modeled Sulphur Variogram Considered for All Resource Domains Grade Interpolation	14-29

Table 14-16:	Rainy River Project Block Model Parameters	14-30

Table 14-17:	Resource Estimation Parameters for Part of the Main Rainy River Deposit Amenable to Open Pit Mining	14-32

Table 14-18:	Resource Estimation Parameters for the Part of the Main Rainy River Deposit Amenable to Underground Mining	14-33

Table 14-19:	Resource Estimation Parameters for the Intrepid Zone	14-34

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Table 14-20:	Search Neighbourhoods Used for Gold and Silver Estimation in Main Rainy River	14-34

Table 14-21:	Search Neighbourhoods Used for Gold and Silver Estimation in Intrepid Zone	14-35

Table 14-22:	Search Neighbourhoods Used for Calcium	14-36

Table 14-23:	Search Neighbourhoods Used for Sulphur	14-37

Table 14-24:	Search Neighbourhoods Used for Grade Estimation in Zone 34	14-38

Table 14-25:	Search Parameters Used to Code the Measured Blocks	14-40

Table 14-26:	Conceptual Assumptions Considered for Open Pit Resource Reporting	14-41

Table 14-27:	Conceptual Assumptions Considered for Underground Resource Reporting	14-42

Table 14-28:	Consolidated Mineral Resource Statement*, Rainy River Project, Ontario, SRK Consulting (Canada) Inc., November 2, 2013^1,2,3,4,5	14-45

Table 14-29:	Mineral Resources¹ for Zone 34 (Domain 200), Rainy River Project, Ontario, SRK Consulting (Canada) Inc., November 2, 2013	14-46

Table 14-30:	Mineral Resources¹ for the Silver Zone (Domain 901), Rainy River Project, SRK Consulting (Canada) Inc., November 2, 2013	14-46

Table 14-31:	Open Pit Mineral Resources¹, Rainy River Project, Ontario, SRK Consulting (Canada) Inc., November 2, 2013	14-47

Table 14-32:	Underground Mineral Resources¹, Rainy River Project, Ontario, SRK Consulting (Canada) Inc., November 2, 2013	14-47

Table 14-33:	Global Block Model Quantities and Grade Estimates¹ at Various Cut-Off Grades	14-49

Table 14-34:	Block Model Quantities and Grade Estimates¹ at Selective Cut-off Grades - Potential Open Pit Mining Material	14-50

Table 14-35:	Block Model Quantities and Grade Estimates¹ at Selected Cut-off Grades Potential Underground Mining Material	14-50

Table 14-36:	Comparison of the October 2012 and November 2013 Mineral Resource Statements	14-52

Table 15-1:	Pit Optimization Parameters	15-5

Table 15-2:	Mill COG Calculated at Various % Dilution and Gold Prices¹	15-8

Table 15-3:	Detailed Open Pit Mine Design Parameters	15-8

Table 15-4:	Estimation of In-pit Dilution and Mine Recovery	15-21

Table 15-5:	Open Pit Mineral Reserves Statement^1,2	15-21

Table 15-6:	Preliminary Estimation of Operating Costs and Breakeven Cut-off Grade	15-25

Table 15-7:	Updated Estimation of Breakeven Cost and COG	15-26

Table 15-8:	Au eq Calculation Parameters	15-27

Table 15-9:	Dilution Estimate for Hangingwall Slough and Blasting Over-break	15-29

Table 15-10:	Dilution Estimate from Backfill Sources	15-30

Table 15-11:	Underground Mineral Reserves Statement¹	15-30

Table 15-12:	Open Pit and Underground Proven and Probable Mineral Reserves^1,2,3,4,5,6	15-31

Table 16-1:	Recommended Overall Slope Geometry by Sector (AMEC, 2013F)	16-3

Table 16-2:	Waste Mine Rock and Low-Grade Ore Stockpile Design Parameters	16-18

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Table 16-3:	Waste Mine Rock Stockpile Characteristics	16-18

Table 16-4:	Low-Grade Ore Stockpile Characteristics	16-19

Table 16-5:	Overburden Stockpile Design Parameters	16-19

Table 16-6:	Overburden Stockpile Design Summary	16-20

Table 16-7:	Operating Shift Parameters	16-22

Table 16-8:	Equipment and Worker Operating Time	16-23

Table 16-9:	Major Equipment Availability and Utilization	16-24

Table 16-10:	Major Equipment Operator Skill	16-24

Table 16-11:	Drill and Blast Specifications	16-26

Table 16-12:	Truck Speed and Fuel Consumption (Loaded and Empty)	16-27

Table 16-13:	Annual Open Pit Mine Equipment Requirements	16-31

Table 16-14:	Annual Hourly Personnel Requirements	16-34

Table 16-15:	Salaried Open Pit Personnel Requirements	16-35

Table 16-16:	Lateral Development Design Parameters	16-42

Table 16-17:	Vertical Development Design Parameters	16-42

Table 16-18:	Stope Design Parameters	16-44

Table 16-19:	Life of Mine Backfilling Waste Rock	16-53

Table 16-20:	Ground Support Recommendations	16-61

Table 16-21:	Mobile Equipment Requirements - Contract Mining Phase	16-66

Table 16-22:	Underground Development and Production Equipment List	16-67

Table 16-23:	Support Equipment	16-67

Table 16-24:	Total Airflow Requirements	16-74

Table 16-25:	Estimated Heating Consumption	16-77

Table 16-26:	Underground Manpower Loading At Full Production	16-92

Table 16-27:	Open Pit and Underground Mine Production Schedule	16-97

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	18-3

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	18-7

Table 18-5:	Proposed Emergency Generators	18-9

Table 18-6:	Tailings Dam Sizing	18-11

Table 18-7:	Construction Fill Materials (Figures 18-1 and 18-2)	18-16

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Table 18-8:	Summary of Water Management Ponds	18-20

Table 18-9:	Water Availability from the Pinewood River below the McCallum Creek Inflow	18-21

Table 18-10:	Summary of WTP Design Criteria	18-25

Table 18-11:	Expected Production of Sludge at Design Flow	18-26

Table 18-12:	Summary of Proposed Habitat Offset Balance for Schedule 2 Amendment Waterbodies	18-31

Table 18-13:	Summary of Proposed Habitat Offset Balance for Section 35(2) Authorization Waterbodies	18-31

Table 20-1:	Local Aboriginal Groups Engaged as Instructed by MNDM, December 2011	20-3

Table 20-2:	Aboriginal Groups Identified by the Provincial Government to be Consulted or Notified, May 2012	20-4

Table 20-3:	Aboriginal Groups Identified by the Federal Government through the Results of Preliminary Depth of Consultation, September 2012	20-6

Table 20-4:	Summary of Concerns and Proposed Approach to Resolve	20-8

Table 20-5:	Species at Risk Known to Occur in the Rainy River Project Environs	20-22

Table 20-6:	Anticipated Federal Environmental Approvals	20-30

Table 20-7:	Anticipated Provincial Environmental Approvals	20-31

Table 21-1:	Project Capital Cost Summary	21-3

Table 21-2:	Direct Costs by Discipline	21-4

Table 21-3:	Civil/Earthwork Quantities	21-5

Table 21-4:	Concrete Quantities	21-5

Table 21-5:	Structural Quantities	21-5

Table 21-6:	Architectural Quantities	21-6

Table 21-7:	Piping Quantities	21-7

Table 21-8:	Electrical Quantities	21-7

Table 21-9:	Labour Productivity Factors	21-10

Table 21-10:	Open Pit Mine Equipment Capital Cost	21-13

Table 21-11:	Site Development and Process Facilities Direct Costs	21-14

Table 21-12:	Indirect Costs	21-16

Table 21-13:	Sustaining Open Pit Mine Equipment	21-21

Table 21-14:	Life of Mine Underground Capital Costs	21-23

Table 21-15:	Underground Capital Cost Breakdown	21-24

Table 21-16:	Underground Capital Development Costs	21-24

Table 21-17:	Underground Equipment Purchases - Mobile	21-25

Table 21-18:	Underground Equipment Purchases - Fixed Plant	21-26

Table 21-19:	Contracted Construction Costs	21-27

Table 21-20:	Annual Project Operating Cost Summary ($ /t milled)	21-30

Table 21-21:	Annual Project Operating Cost Summary ($ M)	21-30

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Table 21-22:	Project Cash Cost Summary	21-31

Table 21-23:	Project Peak Personnel	21-33

Table 21-24:	Open-Pit Mining Cost Summary Per Period	21-34

Table 21-25:	Open Pit Mine Operating Cost Breakdown	21-34

Table 21-26:	LOM Underground Operating Costs by Activity	21-37

Table 21-27:	Underground Operating Costs Mine General Area	21-38

Table 21-28:	Processing Operating Cost Breakdown	21-39

Table 21-29:	Average General and Administrative Costs	21-43

Table 22-1:	Financial Model Criteria and Production Summary^1,2	22-3

Table 22-2:	Operating Costs and Cash Costs over the LOM	22-3

Table 22-3:	Capital Costs over the LOM (M $CAD)	22-4

Table 22-4:	Financial Analysis Summary (Pre-tax and After-tax)	22-6

Table 22-5:	Rainy River Financial Model Summary (CAD$)	22-6

Table 22-6:	Sensitivity Results for Metal Price and Exchange Rate Variations (CAD $M)	22-9

Table 22-7:	Sensitivity Results for Metal Price and Exchange Rate Variations (USD $M)	22-9

Table 22-8:	Selected Sensitivities, Pre-Tax NPV at 5% Discount Rate (CAD $M)	22-9

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

Table 25-2:	Comparison of October 2012 and November 2013 Mineral Resource Statements	25-4

Table 25-3:	Open Pit and Underground Proven and Probable Mineral Reserves^1,2,3,4,5,6	25-5

Table 26-1:	Budget for 2014 (US dollars)	26-1

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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
AACE		American Association of Cost Engineers
AAS		Atomic Absorption Spectroscopy
Actlabs		Activation Laboratories
AF		Aggregate Fill
AI		Abrasion Index
ALS		ALS Minerals Laboratories
AMT		Audio Magneto-Telluric
ANFO		Ammonium Nitrate and Fuel Oil
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
C&F		Cut and Fill
CAD		Canadian Dollar
CAF		Cemented Aggregate Fill
CAPEX		Capital Expense Estimate
CDE		Canadian Development Expenses
CDN		CDN Resource Laboratories Ltd.
CEE		Canadian Exploration Expenditures
CHF		Cemented Hydraulic Fill
CICC		Caracle Creek International Consulting Inc.
CIL		Carbon-in-Leach

NI 43-101 Technical Report

Feasibility Study of the Rainy River Project

LIST OF ABBREVIATIONS


CIM		Canadian Institute of Mining, Metallurgy and Petroleum
CIP		Carbon-in-Pulp
CLRA		Construction Labour Relations Association
CM		Construction Management
CMT		Corporate Minimum Tax
COG		Cut off Grade
CRF		Cemented Rock Fill
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

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LIST OF ABBREVIATIONS


g/t		Grams per tonne
Ga		Billion years
gal.		Gallons
G&A		General and Administrative
GA		General Arrangement
GIS		Gas Insulated Switchgear
GRG		Gravity Reconverable Gold
GSI		Geological Strength Index
h		Hour
ha		Hectare
HAZOP		Hazard and Operability
HBED		Hudson’s Bay Exploration and Development
HF		Hydraulic Fill
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
IRA		Inter-ramp Angles
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

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LIST OF ABBREVIATIONS


LG		Lerchs-Grossmann
LH		Longhole
LHD		Longhole Drill
LHD trucks		Load, Haul, Dump trucks
LHOS		Longhole Open Stoping
LIMS		Local Information Management System
LOA		Living out Allowance
LOM		Life-of-Mine
M		Million
m/h		Metres/hour
Ma		Million years
masl		Meters above sea level
MFL		Manpower Forecasting and Leveling
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
MRC		Mechanized Raise Climber
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

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LIST OF ABBREVIATIONS


NAG		Non-Acid Generating
Nb		Number
NECA		National Electrical Contractors Association
NI		National Instrument
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)
NSR		Net Smelter Return
OB		Overburden
OIQ		Ordre des ingénieurs du Québec
OGS		Ontario Geological Survey
OMC		Orway Mineral Consultants
OMT		Ontario Mining Tax
OP		Open pit
OPA		Official Planning Amendment
OPEX		Operational Expenditure
OSA		Overall Slope Angles
Owner		Rainy River Resources Ltd.
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
PF		Paste Fill
PGA		Peak Ground Acceleration

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LIST OF ABBREVIATIONS


pH		percentage hydrogen
PLC		Programmable Logic Controller
POF		Probability of Failure
POV		Pre-Operation Verification
ppb		part per billion
ppm		part per million
PRV		Pressure Reducing Valve
Q1		First Quarter
Q2		Second Quarter
Q3		Third Quarter
Q4		Fourth Quarter
QA/QC		Quality Assurance/Quality Control
QP		Qualified Person
RF		Unconsolidated Rock Fill
RLOM		Remaining-Life-of-Mine
RPM		Revolutions Per Minute
RQD		Rock Quality Designation
RRGB		Rainy River Green Belt
RRR		Rainy River Resources Ltd.
RRU		Rainy River Underground
RWI		Bond Rod Mill Work Index
S		South
SAG		Semi-Autogenous Grinding
SCIM		Squirrel Cage Induction Motor
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

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LIST OF ABBREVIATIONS


TMA		Tailings Management Area
TMAP		Tailings Management Area Pond
TSX		Toronto Stock Exchange
UCS		Uniaxial Compressive Strength
UG		Underground
USD		United States Dollar
UTM		Universal Transverse Mercator
VCR		Vacuum Recloser
VFD		Variable Frequency Drive
VMS		Volcanogenic Massive Sulphide
VOD		Ventilation on Demand
VOIP		Voice Over Internet Protocol
VPSA		Vacuum Pressure Swing Adsorption
W		West
WBS		Work Breakdown Structure
WCP		West Creek Pond
WDP		Water Discharge Pond
WMP		Water Management Pond
XRD		X-ray diffraction
X		X coordinate (E-W)
mm		Microns
Y		Y coordinate (N-S)
Z		Z coordinate (depth or elevation)
ZBLA		Zoning by-law Amendment

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1.	EXECUTIVE SUMMARY

1.1

Introduction

The Rainy River Project (the “Project”) is an advanced gold exploration project located 50 km northwest of Fort Frances, in northwestern Ontario. A takeover bid for Rainy River Resources Ltd. (“Rainy River” or “RRR”) by New Gold Inc. (“New Gold”) commenced on June 18th, 2013 and following completion of the bid and subsequent compulsory acquisition resulted in New Gold acquiring ownership of 100% of Rainy River’s outstanding shares. The compulsory acquisition was completed as of October 15th, 2013. Rainy River continues to exist as a separate legal entity from New Gold and is a wholly-owned subsidiary of New Gold.

In August 2013, New Gold requested that BBA Inc. (“BBA”) prepare an updated Feasibility Study (the “Feasibility Study” or “the Study”) from the original 2013 Feasibility Study issued on May 23, 2013. This Feasibility Study was prepared and compiled by BBA in a collaborative effort with New Gold and 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”) and to support a feasibility study on the Project as disclosed in New Golds press release entitled “New Gold Announces Its Rainy River Feasibility Study Results”, dated January 16, 2014.

The main intent of the updated Feasibility Study was to ensure that the key inputs and assumptions used for the Project were consistent with those used for New Gold’s other projects and operations as well as provide a technical and economic review of the potential mining operations, based on the Company’s most recent mineral resource estimate (November 2, 2013) and other recent project definition activities such as metallurgical test work. The mineral resources used in the Feasibility Study are based on updated drilling database and block model information as of August 16, 2013 (including the Intrepid Zone). The Feasibility Study is based on the development of a combined open pit (19,500 tpd) and underground mining operation (1,500 tpd), feeding a conventional 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”) 2013 dollars.

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1.2

Contributors and Qualified Persons

This Technical Report has been prepared for New Gold based on work prepared by a number of independent consultants. A summary of the Qualified Persons and Study contributors, their areas 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)	Last Site Visit

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, production scheduling, 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

AMC Mining Consultants (Canada), Ltd.	Underground mine design, production scheduling, capital cost and operating costs.	Colm Keogh Mo Molavi	September 26, 2013 No site visit

AMEC Environment & Infrastructure	Tailings, waste rock and water management, environmental baseline, closure plan. Open pit and underground geomechanics and geotechnical investigations.	David Ritchie Sheila Daniel Adam Coulson	September 4, 2013 May 19, 2011 September 25-27, 2013

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.

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:

•

Direct processing Proven and Probable Mineral Reserves for the combined open pit and underground operations total 66.9 Mt grading 1.54 g/t Au and 3.26 g/t Ag. Total stockpile Proven and Probable Mineral Reserves total 37.4 Mt grading 0.38 g/t Au and 1.99 g/t Ag;

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•

Planned processing rate is 21,000 tpd, and a mine life of 14 years (not including two (2) years of pre-production) or 16 years including pre-production years;

•

Commissioning starts in Q4 2016. Full production will be reached by the end of Q1 2017;

•

Planned overburden removal of 74 M tonnes over a period of seven (7) years;

•

The operational open pit stripping ratio, excluding waste and overburden stripping during the development phase is 3.5:1.0;

•

The overall open pit stripping ratio (LOM) is 3.91 (waste and overburden to ore ratio); and

•

Low-grade ore stockpiles created during the first 8 years will be processed in Years 9 to 14;

Table 1-2: Key Outcomes - Gold Production¹


Parameter	Units	Outcome
Total Average Gold Grade (Years 1 – 9)	g/t		1.44	ILLEGIBLE
Total Average Gold Grade (LOM)	g/t		1.12
Average Open Pit Gold Grade (LOM)	g/t		0.96
Average Underground Gold Grade (LOM)	g/t		4.96
Total Gold Production from Open Pit (LOM)	koz.		2,797
Total Gold Production from Underground (LOM)	koz.		605
Total Gold Production (LOM)	koz.		3,402
Average Annual Gold Production (LOM)	koz.		243
Average Process Plant Gold Recovery (LOM)	%		90.6

1. Total

gold production after mining and processing. Table excludes 0.4 Mt of material milled in the preproduction period.

Table 1-3: Key Outcomes - Silver Production¹


Parameter	Units	Outcome
Total Average Silver Grade (Years 1 – 9)	g/t		3.07	ILLEGIBLE
Total Average Silver Grade (LOM)	g/t		2.80
Average Open Pit Silver Grade (LOM)	g/t		2.48
Average Underground Silver Grade (LOM)	g/t		10.31
Total Silver Production from Open Pit (LOM)	koz.		5,115
Total Silver Production from Underground (LOM)	koz.		889
Total Silver Production (LOM)	koz.		6,004
Average Annual Silver Production (LOM)	koz.		429
Average Process Plant Silver Recovery (LOM)	%		64.1

1. Total

silver production after mining and processing. Table excludes 0.4 Mt of material milled in the preproduction period.

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


Parameter	Units	Outcome
Open Pit Initial Project Capital (incl. Process Plant and Infrastructure)	$M		931	ILLEGIBLE
Total Sustaining Capital³	$M		366
Total Royalty Payments	$M		43
LOM Average Total Cash Cost²	USD/oz. Au		663
LOM Average All-in Sustaining Cost^2,4	USD/oz. Au		765
Pre-tax Net Present Value @ 0% discount rate	$M		983
Pre-tax Net Present Value @ 5% discount rate	$M		462
Pre-tax Internal Rate of Return	%		13.1
Pre-tax Payback Period	Years		5.4
After-tax Net Present Value @ 0% discount rate	$M		774
After-tax Net Present Value @ 5% discount rate	$M		330
After-tax Internal Rate of Return	%		11.3
After-tax Payback Period	Years		5.5

^1.

The financial analysis was conducted using long term consensus averages of USD $1,300/oz. Au and USD $22/oz. Ag. The United States to Canadian dollar exchange rate was assumed to be CAD $0.95:USD $1.00 during pre-production and operations. A discount rate of 5% was used.

²^.

Includes silver credits and royalty payments.

^3.

Funded by internal cash flows.

^4.

All in sustaining costs accounts for operating expenditures and sustaining capital expenditures.

1.4

Property Description

The Rainy River Project is situated in the southern half of Richardson Township, part of the larger Township of Chapple, approximately 50 km northwest of Fort Frances in northwestern Ontario, Canada. The village of Emo is located approximately 25 km to the south on Highway 11. The Project property is approximately 160 km south of Kenora and 420 km west of Thunder Bay.

The Rainy River Project property comprises a portfolio of 162 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 17,018 hectares in the townships of Fleming, Mather, Menary, Pattullo, Potts, Richardson, Senn, Sifton, and Tait. All unpatented mining claims are in good standing and have sufficient work assessment credits available to maintain this good standing for several years. Assessment work of at least $400 per year is required to maintain interest in the staked claims and keep unpatented mining claims valid.

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The Rainy River 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 silts and clays. The Rainy River Project area is sparsely populated.

1.5

Accessibility, Climate, Local Resources, Infrastructure and Physiography

The climate is typically continental, with temperatures ranging from plus 35°C to minus 40°C. Average rainfall in the region is approximately 60 cm, while approximately 150 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.

1.6

Project History

The Rainy River 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 centered 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 that 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 core holes (49,515 m); these were mostly in the Richardson area. The 34 Zone was further drill-tested between 1999 and 2004.

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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 inputted all 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 holes (59,719 m), drilled during 2008. In early 2010, SRK Consulting (Canada) Inc. (“SRK”) prepared a revised mineral resource statement to incorporate information from an additional 124 core holes (68,453 m), drilled on the Project during 2009.

On February 24, 2011, SRK updated the mineral resource statement to incorporate information from an additional 163 core holes (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 core holes, 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 Rainy River’s November 9, 2011 press release and filed by Rainy River on SEDAR on December 23, 2011. A complete summary of drilling on the Rainy River Project between 1994 and 2013 is presented in Chapter 10, in Table 10-1.

A PEA Update was announced in Rainy River’s August 29, 2012 press release and filed by Rainy River on SEDAR on October 12, 2012. A Feasibility Study technical report for the project was filed by Rainy River on SEDAR on May 17, 2013. The mineral resources used in the original 2013 Feasibility Study are based on updated information issued in a press release by Rainy River on October 10, 2012,

In June 2013, New Gold began acquiring ownership of RRR shares pursuant to a takeover bid. Following the completion of a compulsory acquisition as of October 15, 2013, New Gold owns 100% of Rainy River’s outstanding shares. RRR continues to exist as a separate legal entity from New Gold and is a wholly-owned subsidiary of New Gold.

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NI 43-101 Technical Report

Feasibility Study of the Rainy River Project

This Feasibility Report is based on drilling data for the main Rainy River deposit comprising 1,435 core holes (662,849 m) and additional drilling data from 230 core holes (79,575 metres) for the adjacent Intrepid Zone.

1.7

Geological Setting and Mineralization

The Rainy River 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 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 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 westerly plunge of the gold mineralization is the result of high strain deformation 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 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);

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•

Deformed sulphide-bearing quartz veinlets in dacitic tuffs/breccias hosting enriched silver grades (e.g., Intrepid Zone); and

•

Copper-nickel-platinum group metals mineralization hosted in a younger mafic-ultramafic intrusion (e.g., 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 copper-nickel-platinum 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 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 and the greater area surrounding the known zones of 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 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 drill core. A total of 230 core holes have been completed to delineate the Intrepid Zone over a 410 m strike length, tracing the mineralization down dip to a depth of 450 m below surface. The Intrepid Zone has been included in all aspects of this Study.

1.10

Drilling

Nuinsco and Rainy River drilled a combined total of 1,435 core holes (662,849 m) on the main Rainy River deposit between 1994 and 2012. Prior to 1999, Nuinsco also drilled several reverse circulation holes to sample basal till and bedrock for exploration targeting. Reverse circulation drilling data was not used for resource estimation. An additional 230 core holes (79,575 metres) were drilled in the Intrepid Zone since 1996, with the majority of core holes being drilled by Rainy River between August 2012 and June 2013 (225 core holes: 77,969 metres).

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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 with confidence the geometry and boundaries of the gold mineralization.

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 Nuinsco drill core have been identified and selectively sampled by Rainy River. This additional sampling was incorporated into the drill hole database used for mineral resource estimation.

Rainy River used industry best practices to collect, handle and assay core samples collected during the 2005 to 2012 period. All drilling and sampling were 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 analysis and aqua regia digestion and atomic absorption spectrometry for metal analysis. 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 relied partly 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 drill core 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. from SRK has visited the Project property on numerous occasions since 2008. His last visit to the site was from April 30 to May 2, 2013. SRK was given full access to all relevant project data during the mineral resource estimation process.

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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 has summarized the assay results for the external quality control samples on time series, plots bias charts, quantile-quantile and relative precision plots.

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 Test work

1.13.1

Historical Test work

Initial metallurgical test work was carried out at SGS Canada Inc. (“SGS”) in Lakefield, Ontario from 2008 to 2011 and was the basis for the PEA Update issued publicly on October 12, 2012. The test work included mineralization, comminution, gravity separation, flotation, flotation concentrate leaching, and whole ore leaching. Results from the test work 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 ore 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: five ODM, two 433, one CAP, one HS and one 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;

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•

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 zone 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 433 zone samples); and

•

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

1.13.3

Flowsheet Determination Test work

The majority of flowsheet testwork for the Feasibility Study was conducted at SGS between 2011 and 2013. 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% (1% higher than historical test work) 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% (1% lower than historical test work) 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. As such, this process was not further evaluated.

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.

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1.13.4

Comminution Tests

A larger than average scale comminution test work program, providing a high level of confidence in the result interpretation, was undertaken to determine the sizing of the semi-autogenous (“SAG”) mill and ball mill. The larger than average testwork program provided a high level of confidence in the interpretation of the results. 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 (“SMC”). In addition, seven (7) samples were sent to Starkey and Associates for SAGDesign testing.

Results from the test work indicated that the Rainy River material is considered to be hard and relatively resistant to breakage. Based on the test work 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 test work results.

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

Grinding Circuit Design

Four (4) methods and a consultant were used to size the grinding circuit based on the test work results: Morrell’s Equation, JK SimMet with Bond Equation, JK SimMet with Phantom Cyclone, SAGDesign and Orway Mineral Consultants (“OMC”) as an outside consultant.

All methods yielded similar 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.

Recent test work performed on the Intrepid Zone material indicates that the material is more variable than the main pit; however, it should not impact the grinding circuit design if blended at low percentages.

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1.13.5

Gravity Separation

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

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 test work 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 test work was performed to verify the effect of cyanide concentration, preconditioning, use of oxygen versus air and lead nitrate. The variability leaching test work 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 test work results:


Non-CAP Zones		CAP Zone
Au res = 0.0937 • x_Au^0.4223		Au res = 0.2497 • x_Au^1.015
Ag res = 0.0148 • x_Au² + 0.294 x_Ag		Ag res = 0.0363 • x_Au² + 0.245 x_Ag

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

Test work performed on 30 samples from the Intrepid Zone indicated that this material performed comparably to material from the Non-CAP zones. This material has to be blended due to high silver content.

1.13.7

Cyanide Destruction Test work

The SO₂/air cyanide destruction process was investigated on three (3) composites: initial pit, remaining life-of-mine and Intrepid Zone. 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 for the main pit samples were 4.7, 3.5 and 0.1 g/gCN_WAD for SO₂, lime and copper, respectively.

Cyanide destruction test work was also performed on a sample from the Intrepid Zone. The first test yielded high residual cyanide values, however a repeat of the test showed results in line with the main pit.

1.13.8

Environmental Test work

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 Project open pit, and tailings produced in metallurgical test work. 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 the mine rock indicate 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 mine rock 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 to evaluate the long term metal leaching characteristics of these materials. Preliminary results have identified a potential risk of short term neutral metal leaching (cadmium and zinc) from the tailings. These results suggest that a simple lime treatment system may be required for the Tailings Management Area (“TMA”) discharge to the water management pond during operations, but not post-closure.

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

1.13.9

Test work Interpretation

Results from the SGS test work are the basis for the mineral reserve estimate and Feasibility Study. Based on a trade-off study performed in 2012, 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 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 test work was based on cyanide leaching of the gravity tailings.

The extensive grinding test work campaign has allowed for definition of the overall hardness of each zone and indicated that there are a number of 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 five (5) methods used to size the grinding circuit provides a good level of confidence in the sizing of the SAG and ball mill.

Test work performed on the Intrepid Zone indicates that the material should not impact design if blended at low tonnage rates. Given the high silver grade of this material, it is not recommended to be blended at high tonnage due to downstream effects on carbon-in-pulp (“CIP”) and elution.

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 by an average of 0.6% 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) CIP tanks will be required with a total circuit retention time of 117 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.

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1.14

Mineral Resource Estimate

The mineral resource estimate presented herein represents the ninth resource evaluation for the Rainy River Project since 2008. The resource estimate was completed by Dorota El-Rassi, P.Eng. (APEO #100012348) and Glen Cole, P.Geo. (APGO #1416) from SRK, both independent Qualified Persons as defined by Canadian National Instrument 43-101. The effective date of the Mineral Resource Statement is November 2, 2013.

The Rainy River database comprises data from 1,665 core holes (743 km), drilled by Nuinsco and Rainy River up to August 16, 2013. 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. 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. For the main Rainy River deposit, 13 distinct mineralized zones with further domains within these were constructed and used as hard boundaries for estimating the mineral resource. An additional three mineralized zones were defined for estimating the Intrepid Zone mineral resource.

For geostatistical analysis, variography and grade estimation, raw assay data were composited to 1.5 m lengths. The impact of grade capping was analyzed - and capping levels were adjusted for each resource domain and each metal separately to constrain the influence of extreme high grade composite intervals. Capping was applied to the composited data. An unrotated block model was created to cover the entire extent of the Rainy River deposit area. An independent block model was created for the Intrepid Zone. Block sizes varied from 5 x 5 x 5 m for mineral resources amenable to underground extraction to 10 x 10 x 10 m for mineral resources amenable to open pit extraction.

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 separately using ordinary kriging as the principal estimator in each domain from capped composite data within that domain. Grades in domains 601 to 605 were estimated using

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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 mineralization are amenable for open pit extraction, while other parts of the deposits could be extracted using an underground mining methods. 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 Lerchs-Grossman 3D open pit optimizer to develop conceptual open pit shells to delimit the zones of mineralization reported in the mineral resource statement for the Project.

The block models were also reviewed to determine which portions of the gold mineralization have “reasonable prospects for economic extraction” from an underground mine. The mineral resource for the Rainy River Project is reported at two (2) cut-off grades (Table 1-5). The effective date of the Mineral Resource Statement is November 2, 2013. Open pit mineral resources are reported at a cut-off grade of 0.30 g/t gold, whereas underground mineral resources are reported at a cut-off grade of 2.5 g/t gold.

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Table 1-5 - Consolidated Mineral Resource Statement, Rainy River Gold Project,

SRK Consulting (Canada) Inc., November 2, 2013^1,2,3,4,5


			Grade				Metal
	Quantity		Au		Ag		Au		Ag
Category	‘000t		gpt		gpt		‘000 oz		‘000 oz
Direct Processing Material
Open Pit²
Measured		20,282		1.45		1.93		947		1,261
Indicated		80,411		1.35		2.55		3,486		6,584
O/P Measured & Indicated		100,693		1.37		2.42		4,433		7,846
Inferred		9,388		0.97		2.28		292		687
Underground²
Measured		89		4.95		2.75		14		8
Indicated		5,469		4.53		11.34		796		1,994
Measured & Indicated		5,558		4.53		11.20		810		2,002
Inferred		2,641		4.46		8.30		379		707
Stockpile Material³
Open Pit
Measured		6,294		0.37		1.29		74		262
Indicated		64,816		0.44		2.17		919		4,526
Measured & Indicated		71,110		0.43		2.09		993		4,788
Inferred		8,626		0.37		1.16		102		323
Combined Direct Processing and Stockpile Mineral Resources
Measured		26,665		1.21		1.79		1,035		1,531
Indicated		150,696		1.07		2.70		5,202		13,104
Measured and Indicated		177,361		1.09		2.57		6,236		14,635
Inferred		20,655		1.16		2.58		773		1,717

¹	Mineral resources are reported in relation to conceptual pit shells which are limited to 150 m below sea level and are inclusive of the Intrepid Zone.

Open pit mineral resources are reported at a cut-off grade of 0.30 g/t gold, underground mineral resources are reported at a cut-off grade of 2.50 g/t gold based on a gold price of US$1,400 per ounce, a silver price of US$24.00 per ounce, a foreign exchange rate of C$1.10 to US$1.00, gold recovery of 88% for open pit resources, 90% for underground resources and a silver recovery at 75% for all mineral resources.

³	Direct processing material is defined as all mineralization above a cut-off of 0.45 g/t gold and likely to be mined and processed directly.

⁴

Stockpile material includes all material within conceptual open pit shells above a cut-off of 0.30 g/t gold and below a 0.45 g/t gold cut-off as well as material within the CAP zone (code 500) that is suitable for stockpiling and future processing based on average metallurgical recoveries of 88% gold and 75% silver.

⁵	Qualified Persons – The mineral resource statement was prepared by Dorota El-Rassi, P. Eng. (APEO #100012348) and Glen Cole (APGO #1416) from SRK Consulting (Canada) Inc, both independent “Qualified Persons” as that term is defined in Canadian National Instrument 43-101.

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1.15

Open Pit and Underground Mine Design

The evaluation of the open pit reserves was carried out under the supervision of Patrice Live (BBA) and the underground reserves were evaluated by Colm Keogh (AMC). The open pit reserves were generated from the block model provided by SRK effective November 2^nd, 2013 consisting of block sizes measuring 10 m x 10 m x 10 m. The original 2013 Feasibility Study used smaller sized blocks (previously 5 m x 5 m x 5 m). To better optimize the open pit and underground portions of the mineral resources for mine planning, two (2) further block models, covering the underground portion of the main ODM zone and Intrepid Zone were provided by SRK for the estimation of the underground reserves. The underground block models retained the 5 m x 5 m x 5 m block size used in the previous 2013 Feasibility Study.

This Feasibility Study assumes both open pit and underground mining methods will be used for reserve extraction. The milling throughput rate was established at 21,000 tpd (19,500 tpd from the open pit and 1,500 tpd from underground when full production is achieved). The open pit operations will deliver material to a gyratory crusher for primary size reduction and delivery to the processing plant. Ore from the underground operation will be fed to a portable crushing system for size reduction and tramp metal removal prior to being delivered to the gyratory crusher. New Gold’s mining equipment and personnel is expected to be used for the development of the open pit, as well as for the removal of overburden. In addition, a combination of New Gold personnel and various contractors will be used for the underground development and on-going production activities.

1.15.1

Open Pit Mine

Open Pit Mine Design

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

In order to develop an optimal engineered pit design for the Rainy River deposit, BBA used the Lerchs-Grossman 3D (“LG 3D”) routine in MineSight. The pit optimizer searches for the pit shell with the highest undiscounted cash flow. It uses 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 test work), pit slopes (by AMEC based on geotechnical pit slope study) and constraints. 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 Parameters¹


Type of Activity	Unit	Values
Mining Cost of Rock	$/t mined	1.95	ILLEGIBLE
Mining Cost of Overburden	$/t mined	1.50
Processing Cost	$/t milled	8.65
General and Administration Cost	$/t milled	1.21
Gold Recovery	%	89.9
Gold Recovery for CAP zone	%	74.3
Silver Recovery	%	67.1
Silver Recovery for CAP zone	%	69.5
Gold Selling Price	USD/oz.	Intentionally varied from 300 to 1100
Silver Selling Price	USD/oz.	25
Exchange Rate	CAD/USD	1.05
Overall Pit Slope Angle	degree	Varies from 37 to 53 depending on zone
Overall Overburden Slope Angle	degree	16
Pit Limitation		RRP Property
Depth/elevation constraint	masl	-50

^1.

Pit optimization parameters differ from the final operating cost figures. The pit optimization parameters were taken as the initial iteration to determine the optimized pit shell.

Using the technical and economic parameters described previously, the MineSight LG 3D pit optimizer tool was run to produce a series of pit shells for a series of revenue factors. Once the series of pit shells was generated, the total material moved, total open pit resource and stripping ratios were evaluated to identify the optimum pit shell. The selected optimized mineral reserve pit shell for this Study is based on a gold price of USD $800/oz. The selection methodology was based on the stripping ratio and a mine life of approximately 15 years to maximize the NPV.

The optimum pit shell selected was used as a guide to carry out the detailed mine design using the design parameters discussed in Chapter 15. 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 geotechnical design recommendations provided by AMEC.

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1.15.2

Underground Mine

The underground mine design supports the extraction of 1,500 tpd of ore by longitudinal, longhole open stoping (“LHOS”), a mining technique suitable for the geometry and ground conditions of the Rainy River underground resources. Backfilling with cemented aggregate fill (“CAF”) is a significant aspect of the Project with respect to maximization of both resource recovery and mining productivity.

Ore occurs in sub-vertical horizons in varying widths from approximately 3 m to 20 m. Widths over 15 m are rare, and the weighted average across all zones is approximately 8 m. The ore footwall and hanging wall generally dip at 60 degrees or more, but can flatten locally to as low as 45 degrees in some areas.

Ore-grade mineralization occurs in seven independent areas of differing geometry and grade. The spatial separation defines multiple working areas that support the projected 1,500 tpd production rate, and has afforded the flexibility in the production schedule to recover higher grade material earlier in the life of mine.

Rock mass conditions are projected to be generally very good and little water ingress is anticipated.

The underground Mineral Reserves include material from both longhole stopes and the development within mineralized horizons required to access those stopes.

A 3.5 g/t Au-Eq (gold equivalent) cut-off grade was applied for the purpose of delineating the stoping inventory based on a preliminary estimate of operating costs of $112/t, which implied a breakeven cut-off grade of approximately 3.2 g/t Au-Eq. A lower (1.5 g/t Au-Eq) cutoff grade was applied to development within mineralized horizons on the basis that the mining cost is effectively sunk, and the remaining costs to process this material as mill feed are marginal. A more detailed discussion of cut-off grade is presented in Chapter 15.

All underground Mineral Reserves have been classified within the Probable category, and have been estimated from Measured and Indicated Mineral Resources through an appropriate consideration of practical mining constraints, ore recovery estimates and dilution estimates. Inferred Mineral Resources that are contained within the design mining shapes were assigned zero grade values, consistent with the CIM definition for mineral reserves.

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1.15.3

Open Pit and Underground Reserves

The combined open pit and underground mine mineral reserves are summarized in Table 1-7.

Open pit mine reserves are reported at a COG (cut-off grade) 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. The Open Pit reserves are calculated using a mining dilution of 4% at a grade of 0.21 g/t Au and 1.19 g/t Ag. The Underground mineral reserves incorporate 8.3% dilution overall, projected to come from hanging-wall (hw) rock over-break and sloughing, and also from backfill dilution. Rock over-break is included in the stope volumes such that the over-break grade is not reported separately. Backfill dilution is assumed to carry no metal grades. Mining recoveries of 95% and 96.5% have been assumed for open pit and underground reserves, respectively.

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Table 1-7: Open Pit and Underground Proven and Probable Mineral Reserves^1,2,3,4,5,6


Mineral Reserves Category	Tonnage (‘000 t)		Au Grade (g/t)		Ag Grade (g/t)		Contained Metal Au (koz)		Contained Metal Ag (koz)
Direct Processing Material
Open Pit
Proven		15,839		1.47		2.04		746		1,038
Probable		46,866		1.26		3.05		1,896		4,594

Underground
Proven		—		—		—		—		—
Probable		4,187		4.96		10.31		668		1,388

Total Direct Processing Material		66,892		1.54		3.26		3,311		7,021

Stockpile Material
Open Pit
Proven		6,843		0.38		1.51		84		332
Probable		30,541		0.39		2.10		378		2,058

Total Stockpile Material		37,384		0.38		1.99		462		2,390

Combined Direct Processing and Stockpile Material
Open Pit
Proven		22,681		1.14		1.88		830		1,370
Probable		77,407		0.91		2.67		2,275		6,652

Total		100,088		0.96		2.49		3,105		8,022

Underground
Proven		—		—		—		—		—
Probable		4,187		4.96		10.31		668		1,388

Total		4,187		4.96		10.31		668		1,388

Total Combined
Proven		22,681		1.14		1.88		830		1,370
Probable		81,594		1.12		3.06		2,943		8,040

TOTAL		104,275		1.13		2.81		3,773		9,410

^1.

Open pit mineral reserves have been estimated using an optimized pit shell based on metal prices of USD $800 per ounce gold and USD $25 per ounce silver, a foreign exchange rate of CAD$1.05 to USD$1.00, gold recovery of 89.9% (non-CAP zone) and 74.3% (CAP zone) and a silver recovery of 67.1% (non-CAP zone) and 69.5% (CAP zone). The cut-off grade is based on a gold price of $USD 1,200. Underground reserves have been estimated from mining shapes generated using a cut-off grade of 3.5 g/t gold-equivalent. Development material from stope access drives above a cut-off grade of 1.5 g/t gold-equivalent and is also assumed to be sent to the mill for processing. Underground breakeven cut-off grade calculated at 2.75 g/t gold- equivalent based on metal prices of $USD 1,300 per ounce gold and $USD 22 per ounce silver, a foreign exchange rate of CAD$1.05 to USD$1.00, gold recovery of 95% and a silver recovery of 75%.

^2.

Open pit reserves have been estimated using a dilution of 4% at 0.21 g/t Au and 1.19 g/t Ag. An average dilution of 11.7% for the underground stoping (8.3% total underground, inclusive of development ore), which includes dilution from both overbreak and backfill. Open pit and underground reserves have been estimated using a mining recovery of 95% and 96.5%, respectively.

^3.	Open Pit direct processing material is defined as mineralization likely to be mined and processed directly and above a variable cut-off grade ranging from 0.3-0.7 g/t.

^4.	Stockpile material includes all material within designed open pit between variable cut-offs described above in Note 3, as well as material within the CAP zone (code 500) that is suitable for stockpiling and future processing.

^5.

Mineral Reserves for the open pit are derived from the August 6, 2013 resource model. Models for the underground reserves were derived from the August 2013 and September 2013 models for the main ODM zone and Intrepid Zone, respectively. Models were 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 Canadian National Instrument 43-101. The combined mineral resource statement, including the Intrepid Zone was provided to BBA on November 2, 2013. Rainy River’s exploration program in Richardson Township is being supervised by Mark A. Petersen, (AIPG Certified Professional Geologist #10563), Vice President, Exploration for New Gold and a “Qualified Person” as defined in Canadian National Instrument 43-101. New Gold continues to implement a rigorous QA/QC program to ensure best practices in drill core sampling, analysis and data management.

^6.

Qualified persons - The open pit portion of the mineral reserve statement was prepared under the supervision of Patrice Live (OIQ #38991) of BBA, and the underground portion of the mineral reserve statement was prepared by Colm Keogh, P.Eng. (APEGBC #37433) of AMC Mining Consultants (Canada) Ltd., both independent “Qualified Persons” as that term is defined in Canadian National Instrument 43-101.

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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.

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 (AMEC 2012F, 2012G, 2013A, 2013B, 2013C, 2013D, 2013E, 2013F, 2013v) supervised by Adam Coulson (AMEC) and provided to BBA for open pit design and to AMC for underground mine design. This includes an updated review of the Intrepid Zone and a re-evaluation of the geomechanical mine design in the ODM/17 zones (AMEC 2014A).

During the 2012 AMEC geomechanical drilling campaign, assessment of various rock structural domains was based on the analysis, ten (10) NQ-sized core holes totaling 4,500 metres, with geomechanical logging of oriented core and packer testing down the holes to understand the hydrogeological characteristics. These drill holes were oriented in various azimuths, dipping on average at 65° in order to intersect potential open pit walls, known geological features and the three main underground zones of the ODM/17 (West, Central and East). This investigation work was supported with acoustic televiewer surveys by DGI Geoscience Inc. of ten (10) exploration core holes to confirm the orientation of the major joint sets at depth. Subsequent to this, further work was performed by New Gold with orientation of four (4) core holes for structure evaluation in the Intrepid Zone. These boreholes were further geomechanically logged by AMEC to provide design criteria for the Intrepid Zone (AMEC 2014A).

A total of 176 core samples were collected for laboratory strength testing by AMEC, with 268 test specimens 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 waste rock in excess of 1,000 kg each were sampled and crushed for short and long-term strength testing.

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Based on the field and laboratory investigations, open pit slope stability design criteria for optimized 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, were performed with a combination of numerical stress modeling and recognized empirical design tools.

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 geometry and proximity to the surface. The primary equipment fleet at the peak of operations consists of: three (3) hydraulic shovels (2 x 26 m³, 1 x 29 m³), one (1) wheel loader (18 m³), 22 haul trucks (220 t class), three (3) DTH drills (8.5”) and a fleet of support equipment. Operating bench heights of 10 m have been planned for mining operations. Over the life of the mine, a total of 318.2 Mt of waste rock and 73.6 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 318 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 Tailings Management Area (“TMA”) or blended in cement and used as backfill material for the underground mine.

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1.16.2

Open Pit Production Schedule

The open pit mine schedule is shown in Table 1-8.

Table 1-8: Open Pit Mine Schedule


Year (Period)	OP Stripping Ratio (excluding Stockpile)		OP Stripping Ratio (including Stockpile)		Open Pit Direct
					Processing (Mine to Mill)						Mine to Stockpile						Stockpile to Mill						Waste Rock		Overburden
					Mt		Au (g/t)		Ag (g/t)		Mt		Au (g/t)		Ag (g/t)		Mt		Au (g/t)		Ag (g/t)		Mt		Mt
2015 (Y-2)												0.16		0.71		1.84								8.82		13.40
2016 (Y-1)						0.38		1.50		3.00		0.75		0.45		1.76								12.71		11.88
2017 (Y1)		5.15		6.06		7.54		1.44		2.13		6.83		0.42		1.56								26.70		12.13
2018 (Y2)		6.95		7.51		7.65		1.48		1.95		4.30		0.41		1.46								42.72		10.44
2019 (Y3)		7.29		8.13		7.47		1.37		3.03		6.31		0.37		1.91								45.39		9.01
2020 (Y4)		7.63		8.25		7.43		1.31		2.65		4.61		0.41		1.93								44.90		11.80
2021 (Y5)		7.25		7.95		7.24		1.14		4.59		5.03		0.38		2.94								47.63		4.90
2022 (Y6)		6.23		7.06		7.11		1.17		4.18		5.91		0.35		2.43								44.33
2023 (Y7)		4.20		4.63		7.11		1.21		2.59		3.09		0.34		1.78								29.84
2024 (Y8)		1.77		1.84		7.11		1.30		1.88		0.48		0.30		1.20								12.60
2025 (Y9)		0.36		0.71		3.58		1.41		1.59								3.54		0.42		1.50		2.54
2026 (Y10)																		7.12		0.39		1.45
2027 (Y11)																		7.21		0.37		1.88
2028 (Y12)																		7.52		0.34		2.54
2029 (Y13)																		7.66		0.33		2.11
2030 (Y14)																		4.41		0.55		2.33

Grand Total		3.91		6.85		62.62		1.31		2.79		37.46		0.39		1.99		37.46		0.39		1.99		318.18		73.57

Figure 1-1 shows the various stages of the open pit. Phase I is indicated by the pink outline. Phase II is indicated by the green outline and the final pit is outlined in purple.

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LOGO

Figure 1-1: Phase I, Phase II and Final Pit

1.16.3

Underground Mine Infrastructure

Key mine infrastructure includes a 4 km main access decline from a surface portal located to the east of the open pit, internal production ramps servicing each mining zone, ventilation intake shafts equipped with surface heating plant, ventilation exhaust raises, a backfill delivery raise that terminates at an underground truck loading station, a cement storage and grout mixing facility, two main dewatering stations, an equipment maintenance facility, electrical substations, and other smaller, ancillary installations.

Development dedicated to definition drilling activities has been incorporated into the overall design where required.

Figure 1-2 illustrates the geometry of development infrastructure and mining areas against the backdrop of the open pit.

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LOGO

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

1.16.4

Underground Operations

Longitudinal LHOS proceeds along the strike of the orebody in increments (stopes) of dimensions that consider ground quality, support requirements and other practical considerations.

Stopes that are typically 20 m in length along strike and 20 m high underpin the life of mine plan, associated development requirements and mining costs, and the projected productivity of the underground mine. Actual mining experience, particularly with respect to ground conditions, will test the performance of this basic unit, and both height and length can eventually be optimized.

At the Rainy River Project, LHOS in each zone will proceed in a retreating pattern from the strike extent of ore to a common access point on all levels. Mining proceeds upwards from the lowest level of the zone or from an adopted sill elevation, with backfill providing the working platform for each successive lift.

Once mucking of blasted ore is complete, backfilling commences with the placement of a sufficient volume of CAF to provide stability of the backfill that is re-exposed during the extraction of the next stope in sequence. Once this volume of CAF has been placed, un-cemented waste from underground development or from surface is used to fill the remaining void – see Figure 1-3.

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LOGO

Figure 1-3: Typical Stope Long Section and Cross Sections

Trackless mobile equipment will be used for the majority of mining activities.

Ore handling from the underground workings to surface will be accomplished by a modern fleet of 7 m³ loaders and 45t haulage trucks. Loading will occur in close proximity to the stoping areas and will be hauled directly to a surface coarse ore stockpile adjacent to the portal. Primary crushing will be performed on surface in a dedicated jaw crusher as an essential measure to capture tramp steel before the underground ore stream is rehandled into the gyratory crusher that also services the open pit.

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1.16.5

Underground Development and Production Schedule

Underground mine development will commence in 2017 after production from the open pit has started, leading to the projected recovery of 4.2 Mt of ore grading 4.96 g/t Au and 10.31 g/t Ag over a 12 year LOM.

An extensive underground development program that attains a peak of 560 m/month of jumbo advance is required to develop and maintain access to adequate resources to sustain 1,500 tpd of ore production. The ramp-up period to full production from the underground operation will require five years from the onset of underground mine development in 2017. Full production will be achieved in 2022. Waste generated through infrastructure development will be disposed of in underground stopes whenever operationally practical, however, an estimated excess of 1.80 Mt must be hauled to the surface waste stockpile over the life of mine.

Significant vertical raising is required to establish the primary ventilation circuit, and to provide backfill to the underground mine. Infrastructure development, including vertical raising and the construction of underground installations, will largely be accomplished by contractors.

Table 1-9 presents the annual underground ore production schedule.

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Table 1-9: Underground Production Schedule


	Longhole Stoping						Ore Development						Total
Year	(kt)		Au (g/t)		Ag (g/t)		(kt)		Au (g/t)		Ag (g/t)		(kt)		Au (g/t)		Ag (g/t)
2017 (Year 1)		0		0		0		0		0		0		0		0		0
2018 (Year 2)		0		0		0		17		4.65		24.79		17		4.65		24.79
2019 (Year 3)		106		4.75		31.05		92		4.18		34.58		198		4.48		32.69
2020 (Year 4)		185		5.43		39.63		51		4.05		4.22		237		5.13		31.94
2021 (Year 5)		290		5.87		13.85		135		4.57		3.75		425		5.46		10.65
2022 (Year 6)		373		5.24		6.68		178		5.48		2.77		551		5.31		5.41
2023 (Year 7)		387		5.86		3.78		165		3.77		6.64		552		5.23		4.63
2024 (Year 8)		444		5.56		2.58		109		3.32		17.78		554		5.12		5.58
2025 (Year 9)		461		5.19		7.77		89		3.58		12.35		550		4.93		8.51
2026 (Year 10)		401		4.54		4.55		142		4.65		3.28		543		4.57		4.21
2027 (Year 11)		436		4.35		14.17		8		5.59		10.89		443		4.37		14.12
2028 (Year 12)		119		4.29		19.94		0		0		0		119		4.29		19.94

Total		3,201		5.16		10.52		986		4.33		9.64		4,187		4.96		10.31

1.16.6

Proposed Overall Mine Plan (Open Pit and Underground)

The proposed Project will be a combined open pit (19,500 tpd) / underground (1,500 tpd) operation with approximately 104 Mt of material being processed over the life-of-mine at a nominal mill daily throughput of 21,000 tpd (7.67 Mtpa). The open pit mine production rate is planned to be approximately 21,000 tpd of mill feed material initially (2017). Development of the underground mine will start in 2017. The underground mine will come on line in 2019 and the open pit production rate will gradually decrease to 19,500 tpd when the underground operation reaches full production (2022). The mine schedule contains two (2) years of pre-production and envisions a mine operating life of 14 years, exclusive of the pre-production period.

The open pit mine schedule is based on pushback designs by BBA and has followed recommendations as 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 using methods described in Whittle’s EO report. It was conducted to support the PEA Update, the original 2013 Feasibility Study and this Feasibility Study.

Production is based on an elevated cut-off grade strategy, whereby low-grade material will be stockpiled near the primary crusher for future processing. The objective of this strategy is to obtain a higher average milling grade of the open pit direct processing material 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.

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Figure 1-4 presents the total gold and silver production based on the mill feed source of ore: open pit (“OP”)direct processing, underground (“UG”) direct processing and stockpile reclaim. The gold and silver production are based on the project’s recovery curves.

Over the life-of-mine, 2,814 koz. of gold and 5,139 koz. of silver will be recovered after processing from the open pit operation (average open pit LOM grade, including lower grade stockpile: 0.96 g/t Au and 2.49 g/t Ag) and 605 koz. of gold and 889 koz. of silver will be recovered from the underground operation with an average underground LOM grade: 4.96 g/t Au and 10.31 g/t Ag. Total production (including gold and silver produced during the pre-production period) of 3,419 koz. of gold and 6,028 koz. of silver is expected from the Rainy River Project.

LOGO

Figure 1-4: Combined Gold and Silver Production

The combined open pit and underground mine schedule is presented in Table 1-10.

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Table 1-10: Total Milled from OP and UG


	Total Milled UG and OP
Year	(Mt)		Au (g/t)		Ag (g/t)
2014 (Year - 3)		—		—		—
2015 (Year - 2)		—		—		—
2016 (Year - 1)		0.38		1.50		3.00
2017 (Year 1)		7.54		1.44		2.13
2018 (Year 2)		7.67		1.48		2.00
2019 (Year 3)		7.67		1.45		3.80
2020 (Year 4)		7.66		1.43		3.56
2021 (Year 5)		7.67		1.38		4.92
2022 (Year 6)		7.67		1.46		4.27
2023 (Year 7)		7.66		1.50		2.74
2024 (Year 8)		7.66		1.57		2.15
2025 (Year 9)		7.67		1.21		2.04
2026 (Year 10)		7.66		0.68		1.64
2027 (Year 11)		7.66		0.60		2.59
2028 (Year 12)		7.64		0.40		2.81
2029 (Year 13)		7.66		0.33		2.11
2030 (Year 14)		4.41		0.55		2.33

TOTAL		104.3		1.13		2.81

1.17

Process Plant

The process plant facility is designed to have an availability of 92% and a capacity of 21,000 tpd (7.67 Mtpa). This includes an ultimate contribution of 1,500 tpd from the underground mine. Average head grade for the plant is 1.13 g/t Au while the design grade is 2.5 g/t Au, in order to accommodate periods of higher grade feed in the early years of the mine life. The average silver head grade is 2.81 g/t Ag while the design grade is 6.0 g/t Ag.

Run-of-mine material will be crushed to 165 mm in a 1,372 mm x 1,905 mm (54" x 75") gyratory crusher (596 kW) and then stockpiled. The primary crusher building houses the gyratory crusher and the tail-end of the stockpile feed conveyor.

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The crushed rock will be withdrawn from beneath the stockpile, conveyed and ground to approximately 2,800 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 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 the 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 will be returned to the ball mill pump box.

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 and silver 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 solution from carbon stripping will be cooled in a heat exchanger and discharged into two (2) series of electrowinning cells. The gold and silver will be recovered as 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 the pre-detox tailings thickener. The objective of the pre-detox thickener is to recycle residual cyanide from the CIP tailings. Cyanide bearing water is recovered in the thickener underflow and is used to dilute the leach feed. 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 1-5.

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LOGO

Figure 1-5: Schematic Process Flowsheet

LOGO

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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, a portion of the process water for the mill facility will be made-up using reclaim water from the tailings pond area. Additional make-up water will also come from the mine rock pond.

Overall gold recovery of the proposed circuit will be approximately 90.6% for the LOM (life-of-mine) and silver recovery will be approximately 64.1%. It is estimated that 91 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 (“TMA”), emulsion plant, truck shop and various stockpiles is shown in Figure 1-6. Details of the site are described in the following sections.

Site Access

The Project is very well situated where it can benefit from its proximity to existing infrastructure. 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.

LOGO

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Figure 1-6: 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.

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 58 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 Hydro One line currently 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.

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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 truck shop. Potable water will be obtained from the West Creek Pond.

Fresh water will be obtained from the Water Management 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 stockpiles. An additional 26 boreholes were drilled 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 hydrogeologic 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.

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.

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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 can 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 that allows for maintenance of a permanent water cover for closure. The total capacity for tailings produced over the mine life will be approximately 82 Mm³at 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 dams have relatively flat slopes and are constructed largely using mine waste rock placed by the mine fleet, with the dam core, filter, and drain zones constructed by a qualified earthworks contractor.

1.19

Market Studies and Contracts

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

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Environmental aspects have figured prominently in the development of the preliminary layouts and designs for the Rainy River Project described in this Study. 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 Project is unique in that there are no lakes located within, or adjacent to, the main site. Based on multiple years of aquatics baseline investigations, while limited bait fishing does occur within certain area streams, the area does not support a significant 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.

The Rainy River Project is being reviewed through a coordinated Federal Environmental Assessment (“EA”)/Provincial Individual EA. Draft EA Reports were submitted for review by local Aboriginal Groups, and government and stakeholder review. All comments received have been responded to. A Final EA Report was submitted on December 3, 2013 for Federal conformity review. New Gold was informed on December 11, 2013 that the document had passed the conformity review and the report will be issued for a stakeholder and Aboriginal group review period starting in January 2014.

A small number of Federal environmental approvals are likely necessary, as well as an anticipated requirement for a Schedule 2 listing for mineral waste management. The required supporting document for the Schedule 2 listing is included with the Final EA Report. Draft Fisheries Compensation Plans are also included with the Final EA Report. A number of Provincial environmental approvals will also be needed, focused on water taking, effluent and emission management, and closure planning.

The objective of final reclamation for the Rainy River 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 conceptual closure plan consistent with regulatory requirements was part of the Draft EA Reports issued for comment and is included with the Final EA Report. As much as possible, reclamation will be completed progressively during operations, an industry best management practice. The financial model for the Project includes consideration of these costs.

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1.21

Capital and Operating Costs

1.21.1

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 in an open shop environment. Both hourly labour rates were combined in the 80% : 20% ratio (non-building trade workers : building trade union workers) to obtain an average of $132/hour.

The projected pre-production capital cost for the Project is estimated to be $931M, including a $73M contingency allocation (based on 12% of the direct and indirect costs). Total sustaining capital costs for the open pit mine and process facilities are $117.3M, while underground mine development and sustaining capital costs are estimated to be $110M and $138M, respectively. Any working capital requirements for the period between commissioning and first metal sales are assumed to be covered by New Gold’s operating mines. The Project capital summary is outlined in Table 1-11, as follows:

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Table 1-11: Overall Project Capital Cost Summary


Area Description	Pre-Production Capital Costs ($M)		Sustaining Capital Costs ($M)
Overhead Power Line		10.2				ILLEGIBLE
Highway 600 Realignment		12.4
Open Pit Overburden Pre-Stripping		39.3
Open Pit Mine Equipment		85.1		74.1
Mobile Equipment				5.2
Open Pit Waste Removal		45.1
Underground Mine Development Capital¹				110.5
Underground Mine Sustaining Capital²				138.5
Site Development		117.2		4.0
Process Facilities		312.8
Tailings, Water Management and Treatment		50.0		40.5
Fish Compensation Costs				2.0
Chapple Township Compensation				2.1
Equipment Salvage Value				(60.5	)
Reclamation and Closure Costs				49.9

Direct Costs Subtotal		672.1

Indirect Costs (excluding Owner’s Cost)		106.3
Owner’s Cost		79.7

Indirect Costs Subtotal		186.0

Contingency		73.3

Total Capital		931.4		366.3

^1.

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.

^2.

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. Mining equipment, including 10% down payment for 2017, are assumed to be purchased and not leased.

The average unit rate used for overburden removal is $1.53/t, based on using the open pit mining equipment fleet.

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Owner’s costs were provided by New Gold and include capital and operating costs until commercial production is achieved (60% production over a period of 30 days).

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. Full specifications were issued and firm price quotations were obtained from suppliers for the following major equipment: gyratory crusher, apron feeders, SAG mill, ball mill, scalping screen, pebble crusher. 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 capital costs of a total $249M were provided by AMC and are included in the sustaining capital as the open pit operation will have already commenced.

During life of the operation, an estimated sustaining capital of $40.4M is required for necessary expansions, additions or improvements to the tailings management area. This includes $12.7M for the water treatment plant. 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 New Gold mine equipment.

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1.21.2

Operating Costs

Introduction

For operating cost estimate purposes, the Project has been divided into six (6) areas: open pit mining, underground mining, processing, royalties, general and administrative, and refining. All costs are expressed in Q4 2013 Canadian dollars with no allowance for contingency The cost of electrical power was taken to be $0.065/kWh, which includes a $0.02/kWh reduction for the current Northern Ontario Industrial Rebate Program and other potential future government programs. The price of propane and diesel used in calculations were $0.50/L and $0.95/L, respectively. Overall peak operations personnel on site will consist of approximately 606 persons in Year 5 (2021), including 91 employees in the process plant, 29 employees in Administration (G&A), 318 employees in the open pit mine, and 168 employees in the underground mine. Project operating costs by area are summarized in Table 1-12.

Table 1-12: Key Project Operating Costs


Area	LOM Total (CAD$ M)		CAD $ per tonne milled (LOM)		USD $ per gold ounce produced (LOM)
Open Pit Mining (Waste + Rock + Stockpile)^1,2		958		9.22		373
Underground Mining (Cut & Fill)		377		3.63
Processing		961		9.25		268
General &Administration		160		1.54		45
Refining and Transportation³		14		0.14		4
Royalty Payments		43		0.41		12

Subtotal Costs		2,514		24.19		702

Silver by-Product Sales		-139		-1.34		-39

Total Costs Net Silver		2,375		22.86		663

Sustaining Capital		366		3.53		102

All-in Sustaining Cash Costs		2,741		26.38		765

^1.

Equivalent to $1.99 /tonne mined, including stockpile rehandling.

^2.

Stockpile rehandling costs of $0.50 per tonne milled are included in the open pit mining costs.

^3.

Equivalent to $90.10/tonne mined.

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Open Pit Mining

Open pit mining operating costs were developed for the production period (Years 1-9) 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 the Feasibility Study and from BBA’s recently updated database. The LOM mining cost is $2.02/t mined (including waste, ore, overburden and capitalized pre-production costs). The breakdown of operating costs for the open pit is shown in Table 1-13.

Table 1-13: Open Pit Operating Costs


Opex Summary	Total Mine Operating Costs ($M)		Cost ($) per Tonne Mined¹
Hauling		314.4		0.64
Auxiliary Equipment		185.0		0.38
Blasting		152.5		0.31
Loading		149.2		0.30
Maintenance Personnel		79.1		0.16
Salaried Personnel		60.0		0.12
Drilling		49.2		0.10
Services		5.0		0.01

Total Mine Opex²		994.4		2.02

Stockpile Rehandling Cost		50.4		1.34

^1.

Cost per tonne mined includes mining cost for rock ($2.11/t) and overburden ($1.53/t)

^2.

Total Mine Opex includes capitalized pre-production costs. Mine opex during production years is $2.04/t mined and $1.81/t mined during pre-production years.

Stockpile Rehandling

Stockpile rehandling operating costs were developed for the last five (5) years of production (Years 9-14). The costs use the same basis as the mine operating costs and represent the primary and support fleet, fuel and labour required during the reclamation of the low grade ore stockpile. During the stockpile rehandling period, 7.2 Mt to 7.6 Mt of ore per year at an average rehandling cost of $1.34/tonne mined ($0.5/tonne milled) will be trucked to the primary crusher.

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Underground Mining

The operating costs for the Rainy River Underground operation were estimated based on first principle workups of all key mining activities and using a detailed cost model developed by AMC.

Equipment selection and associated operating costs were based on projections for mine activities and workups of required operating hours, productivity and cycle times (hrs/m and hrs/tonne), availability / utilization, and fuel consumption. The operating costs of all major mobile equipment were estimated using supplier quotes and benchmarked against AMC’s database of costs for similar recent projects. Ancillary equipment utilization and costs were calculated based on projected usage to complete general mine tasks.

The unit costs of all major consumables including explosives, ground support, pipes and ventilation ducting were estimated using supplier quotes. The unit costs of power, propane, cement, and diesel were provided by New Gold.

Mine General costs include mine power, heating, technical services consumables, definition drilling, communications and supplies, personal protective equipment and support equipment costs.

Mine maintenance costs include fixed plant and electrical maintenance, maintenance overheads and shop consumables. The mine general and mine maintenance costs were developed based on industry experience and benchmarking against similar projects in the region.

Manpower requirements and labour rates (including benefits and burdens) were developed in consultation with New Gold.

The LOM average underground operating cost has been estimated to be $90.10 /t ore mined. Table 1-14 provides the breakdown by activity. Estimated Mine General costs are further broken down by activity in Table 1-15.

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Table 1-14: LOM Underground Operating Costs by Activity


Operating Costs	Total Cost ($M)		Total Cost ($/t ore mined)
Labour		164.8		39.36
Mine General		57.0		13.61
LH Stoping		48.8		11.65
Ore Development		35.9		8.57
Waste Development		28.9		6.90
Backfill		27.2		6.49
Mine Maintenance		14.7		3.51

Total		377.2		90.10

Table 1-15: LOM Underground Mine General Cost Breakdown


Mine General	Total Cost ($M)		Total Cost ($/t ore mined)
Power		23.2		5.53
Heating		13.0		3.11
Definition Drilling		8.7		2.08
Support Equipment		6.4		1.52
Mine General Consumables		2.1		0.50
Freight		1.8		0.42
Technical Services Consumables		1.3		0.30
Personnel Safety Equipment		0.6		0.15

Total		57.0		13.61

Process Plant

Process plant operating costs were calculated for 14 years of operation. The operating costs are based on metallurgical test work, the mine plan, New Gold salary compensation/benefit guidelines, and recent supplier quotations. The average life-of-mine processing operating costs were determined to be $9.25 per tonne milled at approximately 21,000 tpd. The yearly tonnages from the mine plan vary and were used to obtain the operating costs. The operating cost includes reagents, consumables, personnel (including contractors), electrical power and propane. The consumables accounted for in the operating costs include spare parts, grinding media and liner and screen components. Gold/silver production and operating costs incurred during Q4 2016 are capitalized under the New Gold Owner’s costs until 60% production is achieved in Q1 2017 (commercial production).

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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 New Gold’s recommendations, similar sized operations, and BBA’s in-house database.

The G&A costs were calculated for 14 years of operation and are estimated to average $1.54 per tonne milled. This cost includes:

•

Human Resources;

•

Site Administration, Management and Insurance;

•

Infrastructure Power;

•

Health and Safety Supplies;

•

First Nations Participation Agreement Payments;

•

Security and Paramedic Services;

•

Environmental Costs;

•

G&A Personnel;

•

Information Technology (“IT”); and

•

Training.

Royalties, Refining and Transport

Annual royalty costs (claim blocks and First Nation payments) were provided by New Gold and are based on the conceptual mine design and production profile, along with the terms of the individual royalty agreements. The total estimated LOM royalty cost is USD$ 11/oz Au. Refining and transportation costs are based on a quotation from a North American gold refinery. The total royalty payments are equivalent to $0.41 /tonne milled.

Cash Costs

Table 1-16 provides a summary of average operating cash costs per ounce of gold over the first 9 years and life-of-mine. The operating cash costs include silver credits and royalties.

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Table 1-16: Total and All-in Sustaining Cash Costs

Area

Units

Processing
(Years 1 - 9)

LOM

Open Pit Mining (Waste + Rock + Stockpile)

USD/oz

389.70

372.79

ILLEGIBLE

Underground Mining (Cut & Fill)

USD/oz

Processing

USD/oz

206.58

268.35

General &Administration

USD/oz

35.78

44.68

Royalty Payments

USD/oz

3.65

4.03

Refining and Transportation

USD/oz

10.29

12.03

Cash Costs

USD/oz

645.99

701.89

Silver by-Product Sales

USD/oz

-32.54

-38.83

Total Cash Costs¹

USD/oz

613.46

663.06

Sustaining Capital

USD/oz

122.55

102.30

All-in Sustaining Cash Costs

USD/oz

736.00

765.36

Includes silver credits and royalty payments

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 1-7. 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 1-7: Annual Operating Cash Costs (USD/oz. Au) with Silver Credit

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The average all-in sustaining cash cost was calculated to be USD $736/oz. Au over the first 9 years and USD $765/oz. Au over the LOM.

1.22

Economic Analysis

A financial analysis for the Rainy River 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 New Gold may decide in the future to finance part of the Project using alternative sources of capital. 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 $1,300/oz. and USD $22/oz., respectively;

•

Cash flows are assumed to occur mid-period.

•

The Project value was determined on a pre-tax and after-tax basis and was discounted to the start of Year -2 (2015) which marks the first year of Project construction;

•

Commercial production will begin in January 2017;

•

During the pre-production period and during operations, the United States to Canadian dollar exchange rate was assumed to be CAD $0.95:USD $1.00.

•

All cost and sales estimates are in constant Q4 2013 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 payments and disbursements incurred prior to the effective date of this Study are considered as sunk costs. Disbursements projected for after the effective date of this Study but before the start of construction are considered to take place in the pre- production period;

•

All values shown are post payment of royalties. Total royalty payments over the LOM are $28.6M NPI/NSR (net profit interest / net smelter return) and $14.5M other; and

•

After tax figures assume a combined income tax rate of 25% with 2.7% corporate minimum tax, a mining tax of 10% of taxable mining profits over $500,000 and an allocation of corporate tax attributes among New Gold’s operations.

•

After-tax results were provided by New Gold. BBA has not verified this work.

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The after-tax economic for the Rainy River Project evaluation considered the following taxes:

•

15% federal income taxes on taxable income

•

10% Ontario provincial income tax on taxable income

•

10% Ontario mining taxes on mining income

•

2.7% Ontario Corporate Minimum Tax

Deductions, allowances, and credits were applied to the taxation model where applicable. The most notable of these deductions are Canadian Exploration Expenditures (“CEE”), Canadian Development Expenses (“CDE”), and capital costs eligible for Class 41 of the capital cost allowance system.

The results of the economic analysis summarized below represent forward–looking information as defined under Canadian securities law. Actual results may differ materially from those expressed or implied by forward-looking information. The reader should refer to the Cautionary Note with respect to Forward Looking Information at the front of this Report for more information regarding forward-looking statements, including material assumptions (in addition to those discussed in this section and elsewhere in the Report) and risks, uncertainties and other factors that could cause actual results to differ material from those expressed or implied in this section (and elsewhere in the Report).

The base case general economic results are summarized in Table 1-17.

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Table 1-17: Financial Analysis Summary⁵


Description	Base Case	Units
Ore Milled (LOM)⁶	Million Tonnes		103.9	ILLEGIBLE
Waste Mined (LOM)	Million Tonnes		318.2
Average Gold Grade (Years 1-9)	g/t Au		1.44
Average Gold Grade (LOM)	g/t Au		1.12
Average Silver Grade (Years 1-9)	g/t Ag		3.07
Average Silver Grade (LOM)	g/t Ag		2.80
Open Pit Mining (waste, ore and stockpiling)¹	$/t milled (LOM)		9.22
Stockpile Rehandling	$/t milled (LOM)		0.50
Underground Mining²	$/t milled (LOM)		3.63
Processing	$/t milled (LOM)		9.25
General and Administration	$/t milled (LOM)		1.54
Refining and Transportation Expenses	$/t milled (LOM)		0.14
Royalties	$/t milled (LOM)		0.41
Gold Recovery	% (LOM)		90.6
Silver Recovery	% (LOM)		64.1
Gold Recovered (LOM)⁶	koz. Au		3,402
Silver Recovered (LOM)⁶	koz. Ag		6,004
Initial Project Capital Cost	$M		931
Total Sustaining Capital Cost (including UG development)	$M		366
Total Cash Cost (Years 1 – 9)	USD/oz.		613
Total Cash Cost³ (LOM)	USD/oz.		663
All-in Sustaining Cost (Years 1 – 9)	USD/oz.		736
All-in Sustaining Cost⁴ (LOM)	USD/oz.		765
Net Present Value (0% disc) (Pre-Tax)	$M		983
Net Present Value (5% disc) (Pre-Tax)	$M		462
Internal Rate of Return (Pre-Tax)	%		13.1
Payback Period (Pre-Tax)	Years		5.4
Net Present Value (0% disc) (After-Tax)	$M		774
Net Present Value (5% disc) (After-Tax)	$M		330
Internal Rate of Return (After-Tax)	%		11.3
Payback Period (After-Tax)	Years		5.5

^1.

Equivalent to $1.99/tonne mined, including stockpile rehandling costs. Stockpile rehandling costs of $0.50 per tonne milled is included in the $9.22/tonne milled open pit mining costs.

^2.

Equivalent to $90.10/tonne mined.

^3.

Operating cash costs calculated using the total LOM operating expenditures include silver credits and royalty payments.

^4.

All in sustaining costs are calculated using the total LOM operating expenditures and the total LOM sustaining capital expenditures. Costs include silver credits and royalty payments.

^5.

All values are in Canadian currency unless otherwise noted.

^6.

Excludes 0.4 Mt of material milled from the preproduction period.

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As part of the financial results, the Life-of-Mine Cash Flow Projection has been calculated and is shown in Figure 1-8. The graph includes cumulative cash flow projections (non-discounted) as well as discounted (5%) cumulative cash flow projections on a pre-tax and after-tax basis.

LOGO

Figure 1-8: Life-of-Mine Cash Flow Projections

A pre-tax sensitivity analysis was also performed to ascertain the impact of changes in metal price, capital costs, operating costs and foreign exchange rates. For the purposes of this sensitivity analysis it has been assumed that the royalties remain constant as per the base case. The results the Net Present Value using a discount rate of 5% are shown in Figure 1-9.

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LOGO

Figure 1-9: Pre-Tax Net Present Value (NPV) Sensitivity Analysis at 5% Discount Rate

The effect on the pre-tax NPV’s are shown in Table 1-18 when the CAPEX, OPEX, gold price, silver prices, exchange rate, mining costs and operating costs vary between -30% to +30%. As illustrated in Figure 1-9, the pre-tax NPV is most sensitive to the gold price and the exchange rate.

Table 1-18: Selected Sensitivities, Pre-Tax NPV at 5% Discount Rate (CAD $M)


	Sensitivities
Description	-30%			-20%			-10%		0%		10%		20%			30%
LOM Capital Expenditures		818			699			581		462		343		225			106
LOM Operating Expenditures		975			804			633		462		291		120			(51	)
Gold Price		(525	)		(196	)		133		462		791		1,120			1,449
Silver Price		434			443			453		462		471		481			490
USD Foreign Exchange Rate		1,477			1,139			800		462		124		(215	)		(553	)
Open Pit Mining Costs		742			649			555		462		369		275			182
Under Ground Mining Costs		594			550			506		462		418		374			330
Processing Costs		656			591			527		462		397		333			268

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Table 1-19 shows the pre-tax and after-tax NPV (in $CAD) and IRR sensitivity to varying metal prices and exchange rates. The results are presented in $USD in Table 1-20.

Table 1-19: Sensitivity Results for Metal Price and Exchange Rate Variations, $CAD


Gold Price (USD/oz)		Silver Price (USD/oz)				NPV, 5% Discount Rate (CAD $M)				IRR (%)				Payback Period (Years)
Gold Price (USD/oz)		Silver Price (USD/oz)		CAD/USD		Pre-Tax		After- Tax		Pre-Tax		After- Tax		Pre-Tax		After- Tax
	1,150		20		0.93		147		107		7.8		7.1		6.8		6.8
	1,300		22		0.95		462		330		13.1		11.3		5.4		5.5
	1,450		24		0.97		763		536		17.6		14.9		4.3		4.4
	1,600		26		1.00		1,013		706		21.1		17.8		3.6		3.8

Table 1-20: Sensitivity Results for Metal Price and Exchange Rate Variations, $USD


Gold Price (USD/oz)		Silver Price (USD/oz)				NPV, 5% Discount Rate (USD $M)				IRR (%)				Payback Period (Years)
Gold Price (USD/oz)		Silver Price (USD/oz)		CAD/USD		Pre-Tax		After- Tax		Pre-Tax		After- Tax		Pre-Tax		After- Tax
	1,150		20		0.93		138		100		7.8		7.1		6.8		6.8
	1,300		22		0.95		438		314		13.1		11.3		5.4		5.5
	1,450		24		0.97		738		520		17.6		14.9		4.3		4.4
	1,600		26		1.00		1,009		706		21.1		17.8		3.6		3.8

1.23

Adjacent Properties

Seven (7) properties in the exploration stage are located adjacent to or near the Rainy River 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. Bayfield Ventures Corp. issued a press release on January 14, 2014: “Bayfield Announces Maiden NI 43-101 Technical Report and Mineral Resource Estimate on Burns Block Rainy River Gold-Silver Project, NW Ontario”. 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 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 Project.

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1.24

Other Relevant Data and Information

Project Development Schedule

A Project Development Schedule has been generated to commence production in Q4 of 2016. Full production is assumed to be achieved in Q1 2017. The schedule includes consideration of early work requirements, various studies, the Environmental Assessment process, engineering, procurement, and construction activities. It covers all required site infrastructure, crushing, mineral resource stockpiling, milling and tailings management. The direct on-site personnel requirement peaks at approximately 447 persons during the construction of the Project. The major Project activity durations and milestones are listed below in Table 1-21

Table 1-21: Rainy River Project Development Activities


Activity	Start Date	Completion Date
Detailed Engineering	Q1 2014	Q2 2015
Environmental Assessment, Receipt of Construction Permits	Q2 2013	Q1 2015
Highway 600 Road and Realignment	Q1 2015	Q2 2015
Power Line Construction and Activation	Q1 2015	Q2 2015
Pre-Stripping and Bench Preparation	Q1 2015	Q4 2016
Construction of plant site & infrastructure	Q2 2015	Q4 2016
Commissioning	Q3 2016	Q4 2016
Production	Q1 2017	Q2 2030

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 New Gold Project team.

1.25

Conclusions

The Feasibility Study indicates that the Rainy River Project, based on the calculated Proven and Probable reserves (shown in Table 1-22) of 104.3 Mt grading 1.13 g/t Au and 2.81 g/t Ag, can support a 19,500 tpd open pit and 1,500 tpd underground mine.

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Table 1-22: Proven and Probable Reserves (November 2, 2013)


Resources Category	Tonnage (‘000)		Au Grade (g/t)		Ag Grade (g/t)		Au (In-Situ ‘000 oz.)		Ag (In-Situ ‘000 oz.)
Proven		22,681		1.14		1.88		830		1,370
Probable		81,594		1.12		3.06		2,943		8,040

TOTAL (Combined)		104,275		1.13		2.81		3,773		9,410

Mineralized material will be sent to a process plant designed to achieve gold and silver recoveries of 90.6% and 64.1%, respectively. It is anticipated that, over a mine life of 14 years, approximately 3,402 koz. of gold and 6,004 koz. of silver will be produced excluding capitalized gold and silver produced during the pre-production period.

The initial capital cost of the Project is estimated to be $931M and the sustaining capital (including the development of the underground mine) is estimated to be $366M. The operating life-of-mine total cost is USD $663/oz. Au, including silver credits and royalty payments. The all in sustaining life-of-mine total cost is USD $765 /oz. Au, including silver credits and royalty payments. The Project NPV (pre-tax) is estimated to be $462M and the Project NPV (after-tax) is estimated to be $330M using a discount rate of 5%. The Project internal rate of return (pre-tax) is estimated at 13.1% and the Project internal rate of return (after-tax) is estimated to be 11.3%. The simple payback period (pre-tax) is 5.4 years and the simple payback period (after-tax) is 5.5 years.

The pre-tax sensitivity analyses indicates that positive Project returns can be achieved over the likely range of variation in gold prices (+ 30%/-14%), silver prices (± 30%), capital costs (± 30%) and operating costs (+ 27%/-30%).

Based on the assumptions made in this analysis, it is BBA’s opinion that the Rainy River Project is sufficiently robust to warrant advancing to the next stage of development, that being the start of construction and continuation of detailed engineering.

1.26

Recommendations and Future Work Program

BBA recommends that New Gold proceed with the detailed engineering phase based on the results of this updated Feasibility Study and the Project risks/opportunities identified. However the decision to proceed with a mining operation on the Rainy River Project is at the discretion of New Gold. The suggested work program includes the following components:

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•

Continuation of open pit and underground mine design optimization;

•

Consideration of an underground mining test program within the Intrepid Zone;

•

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;

•

Continuation of key personnel recruitment;

•

Continuation of coordinated environmental assessment process;

•

Continuation of First Nation, Métis and public consultations;

•

Continuation of hydrogeological studies in specific areas;

•

Secure required permits and authorizations from government and regulatory agencies;

•

Continuation of detailed engineering activities; and

•

Construction of the Project (following appropriate approvals).

The 2014 budgeted costs for initiation of key project activities are estimated at approximately US $ 104 M, as shown in Table 1-23.

Table 1-23: Budget for 2014 (US Dollars)


Activity	Cost (US$ M)
Capitalized exploration and condemnation drilling		14.0
Property, plant and equipment payments		60.0
Detailed engineering and studies		20.0
Permitting and environmental monitoring		10.0

Total		104.0

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2.	INTRODUCTION

The Rainy River 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 that comprise Chapple Township, the municipal organization in the area. The Project is located approximately 162 km south of Kenora and 418 km west of Thunder Bay. In June 2005, Rainy River Resources Ltd. (“Rainy River”) acquired the Project from Nuinsco. On June 18^th, 2013 a takeover bid commenced resulting in New Gold Inc. (“New Gold”) acquiring ownership of 100% of Rainy River’s outstanding shares. The acquisition was completed as of October 15th, 2013.

2.1

Scope of Study

The following Technical Report (the “Report”) presents the results of the Feasibility Update Study for the development of the Rainy River Project. In August 2013, New Gold 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, Rainy River Project Director. As of the date of this Report, New Gold is a Canadian publicly traded company listed on the Toronto Stock Exchange (“TSX”) under the trading symbol NGD, with its head office situated at:

Suite 1800, Two Bentall Centre

555 Burrard Street

Vancouver, BC

Tel: (604) 696-4100

This Report, titled “Feasibility Study of the Rainy River 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.

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

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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, production scheduling, processing plant design, site infrastructure, capital costs, operating costs, financial analysis and overall integration.

AMC Mining Consultants (Canada) Ltd.		Underground mine design, production scheduling, capital costs and operating costs.

AMEC Environment & Infrastructure		Tailings, waste rock and water management, environmental baseline, closure plan. Open pit and underground geomechanics and geotechnical investigations.

2.2

Effective Dates and Declaration

This Report is in support of the New Gold press release, dated January 16, 2014, entitled “New Gold Announces its Rainy River Feasibility Study Results”. This report is considered effective as of January 16, 2014. 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 can be significantly more or less favourable.

This Report may include technical information that 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 New Gold and neither BBA nor any affiliate has acted as Advisor to New Gold, 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.

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This report was prepared as National Instrument 43-101 Technical Report for New Gold Inc. (New Gold) by BBA Inc. (BBA), AMEC Environment and Infrastructure (AMEC), AMC Mining Consultants (Canada) Ltd. (AMC) and SRK Consulting (Canada) Inc. (SRK), collectively the “Report Authors”. The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in the Report Authors’ services, based on i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by New Gold subject to terms and conditions of its respective contracts with the Report Authors. Except for the purposes legislated under Canadian provincial and territorial securities law, any other uses 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 re-addressed NI 43-101 Compliant Technical Report: “Feasibility Study for the Rainy River Gold Project”, for New Gold NI 43-101 Technical Report, prepared by BBA Inc., dated July 31, 2013. This Report is publicly available on SEDAR (www.sedar.com).

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 New Gold personnel;

•

Personal inspection of the Rainy River Project property;

•

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

•

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;

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•

Various reports produced by AMEC concerning environmental considerations for permitting, site hydrology, hydrogeology and geotechnical, tailings management, water treatment plant 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;

•

Internal unpublished reports received from New Gold 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.

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 $0.95 = CAD $1.00 was used; and

•

All cost estimates have a base date of the fourth (“Q4”) quarter of 2013.

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.

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2.5

Site Visit

BBA and Rainy River 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 Project property and to review Project development milestones and planning. Rainy River 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 Project property terrain and potential locations for eventual Project infrastructure, Mr. Garett Macdonald led a visit of the Rainy River Project property using a company vehicle. Mr. Colin Hardie visited the site in June 2011; Ms. Sheila Daniel visited the site in May 2011; Mr. Glen Cole was present at the site from April 30 to May 2, 2013, Mr. Adam Coulson visited the site most recently from September 25-27, 2013 and David Ritchie most recently visited the site on September 4, 2013. Mr. Colm Keogh visited the site on September 26, 2013.

2.6

Acknowledgement

BBA would like to acknowledge the general support provided by New Gold personnel during this assignment. Their collaboration is greatly appreciated. The Project also benefitted from the inputs of Mr. Garett Macdonald, Rainy River Project Director, Mr. Paolo Toscano, Director of Engineering and Mr. Nolan Peterson, Project Engineer. 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 lawyers, Davis LLP of Toronto, for the land tenure information summarized in Section 4. An extract from the land title opinion provided by Davis LLP can 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 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, AMC, AMEC or other New Gold 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 NG (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, SAG Design and Outotec.
14.	Mineral Resource Estimate	SRK	D. El-Rassi
15.	Mineral Reserve Estimates	BBA AMC	P. Live C. Keogh	Underground mineral resources estimate by AMC (Section 15.3)

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Section	Description	Responsibility	Qualified Person	Comments and Exceptions
16.	Mining Methods	BBA AMC AMEC E & I	P. Live C. Keogh M. Molavi A. Coulson	Open pit mining and reserves by BBA, underground mining and reserves by AMC (Section 16.3, 16.4 and 16.5). Geomechanical designs by AMEC (Section 16.2.1 and 16.3.6)
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, water management and fisheries compensation by AMEC (Section 18.8, 18.9 and 18.10) and Highway 600 realignment study conducted by TBT Engineering Consulting Group (part of Section 18.1.1). Geotechnical by AMEC (Section 18.1.3).
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 AMC	C. Hardie C. Keogh	AMEC provided tailings area construction quantities, water treatment plant quantities and site closure plan, Merit Consultants provided construction labour rates and productivity factors, AMC provided CAPEX and OPEX for underground mining (Sections 21.5.5 and 21.6.5). New Gold provided royalty costs.
22.	Economic Analysis	BBA	C. Hardie	AMEC provided closure costs, New Gold calculated Project taxes and after-tax cash flow.
23.	Adjacent Properties	BBA	C. Hardie

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Section	Description	Responsibility	Qualified Person	Comments and Exceptions
24.	Other Relevant Data and Information	BBA	D. Runnels	Schedule developed by BBA, and Merit Consultants. AMEC provided permitting information.
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 Project property with the exception of Dorota El-Rassi and Mo Molavi. The information contained in the certificates outline the sections in this Report for which each of the QPs is responsible.


• 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.

• Colm Keogh, P.Eng.		AMC Mining Consultants (Canada) Ltd.

• Mo Molavi, M.Eng., P.Eng.		AMC Mining Consultants (Canada) 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.

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

John Fletcher, C.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

Paolo Toscano, P.Eng.	Metallurgy & Processing	New Gold Inc.

Nick Kwong, P.Eng.	Mining Engineering	New Gold Inc.

Garett Macdonald, P.Eng.	Project Management	New Gold Inc.

Jay Collins, P.Eng., President of Merit Consultants Inc., and John Fletcher, C.Eng., provided construction labour rates and productivity factors. Wayne Clark, of SanZoe Consulting Inc., provided electricity pricing, consulting for applications to IESO and Hydro One and advice for the high voltage power transmission line. Rob Frenette provided road alignments and capital costs for the East Access Road and Highway 600.

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 New Gold Inc. - Rainy River Project (the “Rainy River Project” or “RRP” or the “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 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 and patented mining rights and surface rights land claims located in Mather, Pattullo, 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 Project is comprised of a portfolio of 243 patented mining rights and surface rights lands, including three (3) leasehold patented mining rights, and unpatented mining claims. The Rainy River Project property is collectively located in Fleming, Mather, Menary, Pattullo, Potts, Richardson, Senn, Sifton, and Tait townships. The Project property covers an aggregate area of 17,018.59 hectares.

In Ontario, Crown lands are available to licensed prospectors for the purposes of mineral exploration. Claim staking is governed by the Mining Act (Ontario) 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.

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

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In June 2005, Rainy River Resources Ltd. (“Rainy River”) completed the acquisition of a 100% interest in the Project from Nuinsco Resources Limited (“Nuinsco”). As of November 22, 2013, the Rainy River Project land package consists of:

•

A total of 162 patented mining rights and surface rights land claims, including three (3) leasehold 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.

The unpatented mining claims were independently verified by BBA on December 19, 2013, through the MNDM website (www.mndmf.gov.on.ca/mines/claimaps_e.asp).

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LOGO

Figure 4-1: Location of Rainy River Project (as of November 22, 2013)

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A plan illustrating the distribution of the claims on the Rainy River Project is shown in Figure 4-1 and a title list is provided in Appendix B (claims as of November 22, 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. SRK 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 November 22, 2013)

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4.2

Underlying Agreements

Rainy River, through direct ownership or option agreement, has a 100% interest in the lands forming the Rainy River 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 to acquire five (5) unpatented mining claims in the Tait Township (Figure 4-2) from Perry Vern English for 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 1% of the royalty for CAD $1,000,000.

On December 16, 2011, Rainy River entered into an option 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, repurchase 1% 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, repurchase 1% of the royalty for CAD $1,000,000.

4.3

Environmental Considerations

The RRP area exhibits variable, gently undulating terrain, and is drained principally by the Pinewood River and its associated minor tributaries. The RRP site is located in a low density rural area within the Township of Chapple (total population of 856 in 2006). There is some limited agriculture focused on cattle and fodder cropping, as well as logging activities in the area. Adjacent areas show mainly second growth poplar-dominated forests and wetlands. From an environmental perspective for northern Ontario, the RRP site is unique in that there are no lakes located within, or adjacent to, the RRP footprint. There are small lakes near the proposed transmission line routing.

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Rainy River initiated environmental studies of the Project area in 2008. Initial work from 2008 to 2010 was completed by Klohn Crippen Berger Ltd. (“KCB”). In 2011, Rainy River commissioned AMEC Environment & Infrastructure (“AMEC”) to conduct further environmental baseline investigations which continued into 2013; as well as anticipated mine environmental assessment and permitting stages. Five (5) years of comprehensive environmental baseline studies of the project and regional area have now been completed including:

•

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;

•

Socio-economics;

•

Heritage resources; and

•

Traditional knowledge and traditional land use.

While Rainy River Resources recognizes that the land on which the Rainy River Project is being planned is heavily impacted by historic and ongoing, farming and logging operations, the company feels that it is important that any oral history be properly documented and respected. Rainy River Resources is continuing to support preparation of Traditional Knowledge / Traditional Land Use studies to assess use of the local area by Aboriginal peoples. Rainy River Resources has taken a very proactive approach, and regular meetings are underway with First Nations and Métis to ensure communities are consulted with in a positive manner.

RRR is an active member of the local community with offices in both Emo and Thunder Bay, Ontario that offer residents easily accessible locations to learn about the RRP. RRR continues to engage and consult with the local communities, including First Nations and Métis community members. Through meetings, site tours and regular communications, RRR strives to ensure engagement with all members of the local communities.

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RRR is actively involving local Aboriginal groups in the project planning and has negotiated agreements with First Nations as well as the Métis 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 effects to Aboriginal or Treaty rights. Key issues and interests raised by Aboriginal groups to date are related to: environmental management; employment and benefits; fisheries and wildlife; project components and mining; TLU and culture; and water resources.

Through advice from the Provincial Crown, Aboriginal groups identified to be consulted regarding the RRP are:

•

Mishkosiminiziibiing (Big Grassy River) First Nation;

•

Anishinaabeg of Naongashiing First Nation (Big Island);

•

Naicatchewenin First Nation;

•

Naotkamegwanning (Whitefish Bay) First Nation;

•

Ojibways of Onigaming First Nation;

•

Rainy River First Nations;

•

Buffalo Point First Nation; and

•

Métis - Rainy River Lake of the Wood Regional Consultation Committee Region #1.

RRR will also continue to consult and involve the Fort Frances Chiefs Secretariat and Pwi-Di-Goo-Zing-Ne-Yaa-Zhing Advisory Services Tribal organizations.

Aboriginal groups identified by the Provincial Crown to be notified regarding the RRP are:

•

Anishinabe of Wauzhushk Onigum First Nation (Rat Portage);

•

Couchiching First Nation;

•

Lac La Croix First Nation;

•

Mitaanjigamiing (Stanjikoming) First Nation;

•

Nigigoonsiminikaaning (Nicickousemenecaning) First Nation;

•

Northwest Angle #33 First Nation;

•

Northwest Angle #37 First Nation; and

•

Seine River First Nation.

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4.4

Environmental Approvals in Ontario

The Rainy River 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 Project will require completion of a Federal Environmental Assessment, pursuant to the Canadian Environmental Assessment Act, 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. 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 Provincially-approved Amended Terms of Reference. A Final Environmental Assessment Report is currently under Federal and Provincial review.

After the Environmental Assessment process is completed, environmental approvals will be required to construct, operate and close the Rainy River Project. Due to the complexity and size of the project, various Federal and Provincial agencies will have jurisdiction to provide approvals that will enable project construction to proceed. A schedule has been developed to allow submission of approval applications in parallel with the Environmental Assessment process, such that construction will be able to start once first approvals are received shortly after Environmental Assessment sign-off is received. The Environmental Assessment process is expected to be completed in 2014 with the issuance of various environmental approvals over the ensuing approximate 12 months.

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

The Rainy River 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 F).

5.1

Accessibility

The Project is centred in Richardson Township (part of Chapple Township) in northwestern Ontario, 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 that are of exploration interest. Figure 4-2 in Chapter 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 Project are:

•

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

•

Rainy River, population 909 – 73 km (67-minute drive); and

•

Fort Frances, with a population of 8,103 – 63 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.

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.

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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 150 cm snowfall is recorded annually in the region.

5.4

Physiography

The Rainy River 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 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 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).

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LOGO

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

A	Landscape east of the open pit area, looking north towards the Intrepid Zone within the Project Area.

B	View of the Rainy River Exploration facility. A new facility was constructed in 2013 a few kilometres east of the Project Area.

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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 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 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 Project region. There was insufficient encouragement to continue and exploration was curtailed.

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.

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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 km^2
“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.

1995

Discovery of 34 Zone (copper-nickel-platinum group metals) followed by intensive diamond drilling.

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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

Eight (8) previous mineral resource evaluations were prepared for the Rainy River Project by Mackie et al. in 2003, Caracle Creek International Consulting Inc. (“CCIC”) in 2008 and by SRK in 2009, 2010, 2011 and 2012. Six (6) 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 Statement¹ for the Rainy River 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

^1.	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 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.

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Table 6-3: Mineral Resource Statement¹ for the Rainy River 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 Pit²
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
Underground²
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

^1.

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.

^2.

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 Statement¹, Rainy River 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 Pit²
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
Underground²
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

^1.

^2.

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.

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.

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Table 6-5: Mineral Resource Statement¹, Rainy River 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 Pit²
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
Underground²
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

^1.

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.

^2.

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.

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.

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Table 6-6: Mineral Resource Statement¹, Rainy River 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 Pit²
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
Underground²
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

^1.

^2.

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.

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.

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Table 6-7: Mineral Resource Statement¹, Rainy River 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 Pit²
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
Underground²
Measured³		89		4.62		2.55		13		7
Indicated³		3,083		4.32		5.00		429		495
Measured and Indicated³		3,172		4.33		4.93		442		502
Inferred³		1,172		4.12		5.82		155		219
Combined Mining
Measured³		23,243		1.30		2.00		973		1,498
Indicated³		127,337		1.14		2.54		4,690		10,410
Measured and Indicated³		150,580		1.17		2.46		5,663		11,908
Inferred³		88,288		0.78		2.35		2,217		6,678

^1.

^2.

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.

^3.	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.

In October 2012, SRK prepared an eighth mineral resource statement to consider new drilling data available up to July 10, 2012. The total database comprised 1,435 core boreholes (662,849 m). This consolidated mineral resource statement dated October 10, 2012 is represented in Table 6-8.

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Table 6-8: Consolidated Mineral Resource Statement¹, Rainy River Project, Ontario

SRK Consulting (Canada) Inc., 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
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
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		2.46		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

^1.

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.

^2.**

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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7.	GEOLOGICAL SETTING AND MINERALIZATION

7.1

Regional Geology

The Rainy River 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 of the area west of Fort Frances is shown in Figure 7-1. In the region east of Fort Frances, the Wabigoon subprovince is bound 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 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.

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.

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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 Project.

Three (3) phases of the Quaternary Wisconsinan glaciation are recorded in the Rainy River 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 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, electrically 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)

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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 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 Project area was therefore successively covered by the Labradorean and Keewatin ice sheets.

7.2

Property Geology

The geology of the Rainy River Project is inferred 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 Project is shown in Figure 7-2.

The Rainy River 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 Project (from Rainy River, 2013)

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7.2.1

Lithology

Lower Mafic Volcanic Succession

The Rainy River 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 noted interbedded or intruding respectively 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.

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.

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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

The volcano-sedimentary sequences of the Rainy River Project and regional greenstone belt were affected by at least five (5) main deformation episodes. See Figure 7-3.

All rock types within the Intrepid mineralized zone, with the exception of the late diabase dykes, have a well-developed, moderately south-dipping, penetrative foliation and a moderately southwest-plunging stretching lineation (Figure 7-4).

These fabrics commonly obliterate the original layering and a primary folding event (F1), consisting of large recumbent and low-plunging folds with a north-south axial plane parallel to the strike of a steep axial-plane foliation (S1). The folding event was also accompanied by localized T1 thrusts (potentially at regional scale) (Rankin 2013). The pre-D1 mineralized veins are strongly folded and commonly transposed into the S1 foliation.

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Figure 7-3: Regional Structural Trends on the Rainy River Gold Project, Interpreted from Aeromagnetics (Rankin, 2013)

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A second folding event (F2) consisting of east-southeast-trending upright to overturned shallow-plunging folds of variable intensity, refolded S1 and L1 with variable dips and plunges across the belt. The Rainy River Project auriferous zones lie within a moderate to steeply dipping F2 limb with So & S1 trending 110/55 (average) and L1 steep south-southwestern plunge (~down-dip). This forms the northern limb of a regional F2. F2 folding was possibly accompanied by T2 thrust to high angle reverse faults, partitioning subdomains with varying D2 strain. A weak F2 axial-planar foliation is locally visible in both drill core and outcrop. F2 fold axes are typically subhorizontal to shallowly plunging (Rankin 2013). Steep-plunging ore-shoots within the mineralized zones probably represent localized F1 fold hinges, forming thickened zones of early veins. Termination of ore shoots down-plunge may locally be due to refolding of F1 about local F2 folds at an oblique angle.

Broad-scale bends (D3) in the D1/D2 structural grain followed as kink folds in the greenstone belt. These trend north-northwest- to north-east (with some conjugate kink geometry evident). F3 folds are associated with subvertical S3 spaced fracture cleavages to small scale faults. See Figure 7-5.

A consistent sinistral displacement along these structures may be due to progressive rotation of the compressive stress direction from D3 to D4. Small-scale F3 kinks are common within the layered sequences in outcrop and drill core. Very localized remobilization of quartz-sulphide (as veinets) into the kink axial planes may have produced small zones of enriched mineralization. F3 fold axes are typically steeply plunging (where is folding steeply dipping So/S1 fabrics). Emplacement of ~north-trending granitoid stocks east of the Rainy River gold project is interpreted to have occurred along F3 kink axes (possible reactivated basement faults) (Rankin 2013).

D4 is represented by a late-stage NNW-SSE to N-S compressive episode causing broad warping of all pre-existing fabrics, including F3 mega-kink axial planes. D4 is interpreted to have also caused both flat-lying breccia bodies with late-stage kaolin-sericite alteration in the Intrepid area (subhorizontal tension gash structures), and a weak ESE-trending foliation in the Blackhawk granitoid stock.

The final deformation event; D5 is represented by late stage (Proterozoic) emplacement of NW-trending mafic dykes (NE-SW extension).

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Figure 7-4: 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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Figure 7-5: 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.

Outcrop (426255E/5409525N) showing set of 360° to 020°-striking, subvertical faults offsetting mafic dyke predominantly with sinistral separation in this area.

A late-stage north-northwest to south-southeast to north-south compressive episode caused broad warping of all pre-existing fabrics, including F3 mega-kink axial planes (Figure 7-6). This D4 episode is interpreted to have also caused both flat-lying breccia bodies with late-stage kaolin-sericite alteration in the Intrepid Zone (subhorizontal tension gash structures) and a weak east-southeast trending foliation in the Blackhawk granitoid stock.

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Figure 7-6: Pressure Shadows Around Rigid Objects in Dacitic Rock

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 7-7); and

•

Later quartz-sulphide veins are variably deformed and overlap in time with the main regional deformation.

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

This strongly suggests that the current geometry and plunge of the gold mineralization at the Rainy River Project is the result of high strain deforming features associated with gold mineralization and rotating the ore plunge parallel to the stretching direction (Figure 7-8).

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Figure 7-8: Structural Control of the Plunge of the Gold Mineralization at the Rainy River 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 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 (Intrepid Zone); 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 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 7-9). High-grade gold mineralization in this zone is associated with deformed quartz-pyrite-gold veinlets (Figure 7-10) 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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Figure 7-9: 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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Figure 7-10: 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).

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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 7-11).

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Figure 7-11: 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 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.

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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.

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. See Figure 7-12. Relative to ODM/17 and 433 Zones, the CAP Zone has a higher pyrite-chalcopyrite content.

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Figure 7-12: Higher Grade Gold Mineralization—Borehole NR10-474 from 188.0 to 234.0 m. (SRK, 2011)

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.

High-grade gold and silver mineralization in the Intrepid Zone is associated with deformed quartz-pyrite-gold, quartz-pyrite-silver, or quartz-pyrite-gold-silver veinlets that overprint other mineralization styles. See Figure 7-13. The gold-silver ratio is determined by their location within the base metal zonation controlling the low to moderate grade mineralization.

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Figure 7-13: Intrepid Zone Gold Mineralization. Deformed Pyrite-Sphalerite Veins and Stringers Borehole NR131542 (Rainy River, 2013)

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

The origin of gold mineralization at the Rainy River 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 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) suggest 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 Project is associated with strong sodium depletion, potassium enrichment, aluminous alteration, a strong gold-pyrite association, common 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.

Siddorn (2007) and Hrabi and Vos (2010) suggested that at least two (2) stages of gold mineralization occurred in the Rainy River 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).

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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)

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).

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.

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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 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 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 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 Project also contains nickel, copper and platinum group metals sulphide mineralization associated with a differentiated ultramafic-mafic intrusion. This 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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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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Figure 8-3: Schematic Section and Hydrothermal Alteration Associated with the

Rainy River 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 Project property during 1990 to 2004 is discussed in Section 6.1.2 (refer to Appendix C).

9.3

Rainy River Exploration Work (2005-2013)

In June 2005, Rainy River completed the acquisition of a 100% interest in the Project from Nuinsco.

In the same year, Rainy River re-logged 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 Project. In 2008, Rainy River drilled an additional 100 core boreholes (53,123 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 the core boreholes drilled during 2008.

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.

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In early 2011, SRK updated the mineral resource statement to incorporate information from 165 core boreholes (84,134 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 additional core boreholes completed on the project between December 2010 and February 2011. This mineral resource statement represented the sixth resource evaluation prepared for the Rainy River 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. See 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.

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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 388 core boreholes (188,588 m) considering data up to January 9, 2012. This mineral resource statement represents the seventh resource evaluation prepared for the Rainy River Project.

In August 2012, exploration drilling discovered a significant new gold and silver zone approximately 1 kilometre east of the proposed open pit boundary of the Rainy River Project. See Figure 9-2. Borehole NR121258 intersected 2.2 gpt gold and 38.5 gpt silver over 18.5 metres, including 6.0 gpt gold and 83.9 gpt silver over 3.0 metres at a vertical depth of 210 metres. This new zone, termed the Intrepid Zone, clearly demonstrates the potential for new discoveries along strike of known mineralization.

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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 represented the eighth resource evaluation prepared for the Rainy River Project.

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Figure 9-2: Section through the Intrepid Zone

The Intrepid Zone was covered by a mobile metal ion (MMI) soil survey and identified anomalous gold over a prominent magnetic low trend similar to the magnetic trend that hosts the majority of the Rainy River Project deposits. The new zone contains disseminated and fracture-related mineralization, including 2 – 3 percent pyrite and variable amounts of sphalerite, galena, and chalcopyrite. Both electrum and visible gold have also been identified in core. Up to August 16^th, 2013, a total of 230 boreholes (79.575 m) have now penetrated the Intrepid Zone over 410 metres strike length, and have traced the mineralization down dip for 550 metres.

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Wide zones of near-surface gold mineralization continue to contribute to mineralization that has the potential to be exploitable via open pit mining, complemented by higher grade intersections that have the potential to be extracted by underground mining early in the mine life. Diamond drilling in the third and fourth quarters of 2013 has shifted away from resource drilling to test regional targets, infrastructure condemnation and, potential fold repetitions of the Intrepid Zone horizon elsewhere on the Rainy River Project property. The alteration envelope associated with the Intrepid Zone mineralization has been traced by widely spaced boreholes (400 metres) along strike on an interpreted fold limb.

Exploration drilling undertaken in 2013 after the drilling data cut-off of August 16^th, 2013 includes 27 diamond drill holes for a total of 7,960.5 metres. As mentioned above, the majority of this drilling was conducted outside the main mineralized zones. None of the logging or analytical results from this drilling post the data cut-off date indicate any material changes to the results or conclusions contained in this report.

Additional exploration work post August 16^th, 2013 includes a program of 2,085 MMI soil samples in five (5) discrete grids was undertaken in the second half of 2013 to detect geochemical anomalies through the overburden cover to assist in identifying Intrepid-like zones along these structural trends.

Further, a re-logging program was undertaken to update and re-interpret lithological units within the main ODM zone of the Rainy River Property with an ultimate goal of producing a set of improved, geologically consistent sections and plans for this zone. This in turn will help drive future drill planning and ultimately improve confidence in future resource definition. Phase 1 of the ODM re-logging Program was completed in November 2013 and a total of 56,000 m of core was re-logged. As of year-end 2013, sectional and level plan interpretation had begun and was progressing on schedule.

To date, the Rainy River 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 recently discovered Intrepid Zone. The HS, New and 433 Zones are located north of the ODM, while the CAP Zone is to the south.

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The exploration activities undertaken by Rainy River between 2005 and 2013 are summarized in Table 9-1

Table 9-1: Summary of Exploration Work by Rainy River on the

Rainy River Project between 2005 and 2013


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

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Activity	Date	Performed by
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.
2009
Preliminary Pit Slope Design and Waste Management Assessment	Nov 2008-Feb 2009	Klohn Crippen Berger Ltd.
Preliminary Pit Slope Stability Assessment	Nov 2008-March 2009	Klohn Crippen Berger Ltd.
DD 139 Holes (69,753 m) in 17/ODM, 433,	Jan-Dec 2009	W. Rayner, B. Nelson, A. Tims, C. Hercun,
CAP, South and BP Zones		A. Shute
Age Dating of Lithologies	Jan-Feb 2009	University of Toronto Geochronology Lab
Third Mineral Resource Statement	Apr 2009	SRK Consulting (Canada) Inc.
Surficial Drainage Project	Apr-Sept 2009	K. Smart Associates Limited
Socio-Environmental Baseline Assessment.	May-Oct 2009	Klohn Crippen Berger Ltd.
Phase 5, 28-Hole RC Program	Jun-Aug 2009	Overburden Drilling Management
Acid Leach Test	Jun-Aug 2009	Klohn Crippen Berger Ltd.
New Office Building	Jul-Oct 2009	W. Rayner, K. Schram, B. Burnell, R. Burnell
Lidar Survey	Sept 2009	Lidar Services International
Preliminary Metallurgical Testing	Sep-Nov 2010	SGS Canada Inc.
2010
DD 196 holes (88,61 m) in 17/ODM, 433,	Jan-Dec 2010	W. Rayner, B. Nelson, A. Tims, C. Hercun,
CAP, South and BP Zones		A. Shute, J. Pattison, H. Buck, K. Pedersen
Phase 6, 37-hole RC Program (1,066 m)	Jan-Feb 2010	Overburden Drilling Management
Metallurgical Testwork	Jan-Dec 2010	SGS Canada Inc.
Environmental Baseline Studies	Jan-Dec 2010	Klohn Crippen Berger
DD 4 Geotechnical Drill Holes (1,405 m)	Mar-May 2010	Klohn Crippen Berger Ltd.
Review of Pit Slope Design	Mar-May 2010	SRK Consulting (Canada) Inc.
Memorandum of Understanding with Fort Francis Chiefs Secretariat	May 2010	Rainy River Resources Ltd.
Structural Study	May-Sept 2010	SRK Consulting (Canada) Inc.

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Activity	Date	Performed by
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
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
Infill Sampling of Historical Drill Holes - 9504 Samples (13,660 m)	Jul-Aug 2010	A. Tims, C. Hercun, A. Shute, H. Buck, J. Pattison
New Core Logging Facility	Jul-Dec 2010	C. Hercun, True-line Construction
Phase 7, 34-Hole RC Program (762 m)	Sept-Nov 2010	Overburden Drilling Management
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
2011
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
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.
Fifth Mineral Resource Statement	April 2011	SRK Consulting (Canada) Inc.
Environmental Baseline Gap Analysis	Apr-May 2011	AMEC Earth and Environmental
First Quarter QA/QC Report	Apr-June 2011	Analytical Solutions Ltd.
Fugro AEM Survey	May 2011	Fugro Airborne Surveys Corp.
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
Preliminary Economic Assessment of the Rainy River Gold Property	December 2011	AMEC, BBA Inc., SRK Consulting, Golder Associates
2012
2011 Rainy River Gold QC Report	Jan 2012	Analytical Solutions Ltd.
DD 237 Holes (95,760 m)	Jan-Jul 2012	Rainy River Resources Ltd.
DD 403 boreholes (141,987 m) in 17/ODM, 433, The Gap, South, Beaver Pond, District, Northern, and HS zones plus infrastructure condemnation drilling	Jan-Sept 2012	A. Tims, C. Hercun, A. Shute, H. Buck, J. Pattison, K. Pedersen, A. Philippe, R. Montufar, M. Rousseau, C.Shaw
Seventh Mineral Resource Statement	Feb 2012	SRK Consulting (Canada) Inc.
Mobile metal ion soil surveys - various	May-Oct 2012	Rainy River Resources Ltd.
Intrepid DD 102 boreholes (28,689 m)	May, July-Dec 2012	A. Tims, A. Shute, H. Buck, J. Pattison, M. Rousseau, C.Shaw, D.Hyde

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Activity	Date	Performed by
Report on 34 zone & Pinewood Ni, Cu & PGE mineralization	July 2012	Revelation Geoscience Ltd.
Eighth Mineral Resource Statement	Oct 2012	SRK Consulting (Canada) Inc.
Preliminary Economic Assessment Update of the Rainy River Gold Property	Oct 2012	AMEC, BBA Inc., SRK Consulting, Golder Associates
2013
DDH gyro survey 53 boreholes	Jan-Feb 2013	Reflex Instrument North America
Intrepid specific gravity data	Jan-June 2013	ALS Chemex Laboratory
Intrepid DD 122 boreholes 2 geotechnical, 4 oriented core	Jan-June 2013	A. Tims, A. Shute, H. Buck, J. Pattison, M. Rousseau, C.Shaw, D.Hyde
Soil gas hydrocarbon orientation survey	April-May 2013	Rainy River Resources Ltd.
Feasibility study of the Rainy River Gold Project	May 2013	AMEC, BBA Inc., SRK Consulting, Golder Associates
MSc thesis - style, geometry, timing and structure	Jun-Sept 2013	M. Pelletier - Institut national de la recherche scientifique
Feasibility Study of the Rainy River Gold Project: Re-addressed to New Gold	July 2013	AMEC, BBA Inc., SRK Consulting, Golder Associates
2,085 MMI Sampling program across 5 grids	July-October 2013	Rainy River Geologists
56,000 m re-logging program within ODM	August-November 2013	Rainy River Geologists
Ninth Mineral Resource Statement - reported herein	November 2013	SRK Consulting (Canada) Inc.
Exploration drilling 27 boreholes, 7,960.5 m	August-December 2013	Rainy River Geologists

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10.

DRILLING

10.1

Drilling from 2004 to 2013

10.1.1

Introduction

Nuinsco and Rainy River Resources drilled a combined total of 1,435 core boreholes (662,849 m) on the main Rainy River deposit 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. An additional 230 boreholes (79,575 metres) were drilled in the Intrepid Zone between 1996 and August 16, 2013. The majority of boreholes were drilled by Rainy River between August 2012 and June 2013 (225 boreholes: 77,969 metres). The mineral resources reported herein consider all drilling data available as of August 16, 2013.

The distribution of the core drilling in the area of the main Rainy River deposit is summarized in Table 10-1 illustrated in

Figure 10-1 and Figure 10-2. The Nuinsco data contributes about 7% of the total drilling information (measured by drill metres). Rainy River Resources drilling data represents 93% of the drilling data.

Table 10-1: Core Drilling Completed on the Rainy River Project (1994-2013)¹


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
		1996 - 2013	²		Core		230		79,575

Total		1994 - 2013	³		Core		1,665		742,424

^1.	Table reflects all exploration drilling and will differ from the total meterage incorporated into resource calculations.

^2.	Drilling within the Intrepid Zone.

^3.	2013 represents the August 16^th, 2013 drilling data cut-off date for the mineral resource calculation.

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LOGO

Figure 10-1: Core Drilling Data by Period (1994 to August 16^th, 2013)

LOGO

Figure 10-2: Drill Collar Plan in Relation to Resource Domains within the Main Rainy River Area

and Showing Conceptual Pit Outline

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Exploration drilling undertaken in 2013, after the drilling data cut-off date (August 16^th, 2013), includes 27 diamond drill holes for a total of 7,960.5 metres. The majority of this drilling was conducted outside the main mineralized zones. None of the logging or analytical results from this drilling post data cut-off indicates any material changes to the results or conclusions contained within this report.

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, therefore 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 on the main Rainy River deposit has been completed by Bradley Bros. Ltd, whereas drilling on the Intrepid Zone has been completed by Naicatchewenin Development Corporation (“NDC”) in partnership with C3 Drilling and Major Drilling Group International Inc. 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 using a known control station location.

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 late 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

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successful and implemented for almost a year. At the Intrepid Zone, 60 out of the 230 boreholes have been resurveyed with a Reflex Gyro at 5-metre intervals. The initial orientation of the Gyro instrument was set using an azimuth pointing system (“APS”) at the collar location. In SRK’s opinion, the survey method used by Rainy River is appropriate.

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 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 metres. Core recovery is reported to be excellent, but was not measured until 2008 when procedures were upgraded to include core recovery and geotechnical parameters such as rock quality designation (“RQD”), joint/fracture analyses, material type, and rock strength. Structural and geotechnical logging is usually not based on orientated core. More recently, Rainy River submitted several samples per borehole for specific gravity analyses for both waste rock and mineralized zones.

Core was not routinely photographed, although significant intersections and features are photographically recorded. Diamond core is archived in secure, high quality storage facilities on the Rainy River Project site where it is under security watch 24 hours a day, seven days a week.

Following the drilling data cut-off date for the mineral resource calculation, minor changes to the drilling procedures were implemented to improve data confidence and usability. These include an increase in the sampling density for condemnation holes where these are being selectively sampled. The frequency of systematic insertion of certified reference standards and blanks into the sample batches to be shipped to the lab was increased slightly. The collection of routine geotechnical data such as RQD and the practice of core photography have been extended to all exploration drill holes.

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10.2

Drilling Pattern and Density

Rainy River 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.

The Intrepid Zone boreholes were mostly drilled at a dip of 51 degrees and occasionally up to 61 degrees on north directed azimuths. The zone has been drilled on at least 25 x 25 metres centres; however, there can be significant deviation with depth. Before 2012, drilling was targeting geophysical conductors or was designed as stratigraphic fences. Early 2012 boreholes were spaced 200 metres apart to complete infrastructure condemnation. After the discovery at NR12-1258, the resource drilling began at 50-metre borehole spacing. By January 2013, the drilling pattern was reduced to 25 – 30 metres.

10.3

SRK Comments

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.

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11.

SAMPLE PREPARATION, ANALYSES, AND SECURITY

11.1

Sampling Method and Approach

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 Project property on numerous occasions, and most recently from April 30 to May 2, 2013, drilling and sampling was active. The following information regarding Rainy River sampling procedures (2005 to 2013) 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 ensure 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 Gardewine North Shipping, in Fort Frances. A typical dispatch contains approximately 400-600 samples. Rice bags requiring overnight storage are securely stored inside a designated building.

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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 the generally accepted industry best practices. SRK concludes that the samples are representative of the source materials and there is no evidence of bias.

11.2

Sample Preparation and Analyses

11.2.1

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.

11.2.2

Rainy River Samples

Between 2005 and August 16^th, 2013, Rainy River collected core samples from 1,236 core boreholes (613,498 m) from the main Rainy River deposit and from an additional 221 core boreholes (73,578 m) from the Intrepid Zone. 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.

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

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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 – 2013) is provided herein.

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:

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•

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.

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.

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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.

ALS Minerals (2011 – 2013)

ALS Minerals 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.

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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).

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.

11.2.3

Metallurgical Testing

Rainy River used the SGS Canada Minerals Services Lakefield Laboratory in Lakefield, Ontario (Lakefield) for metallurgical testwork. 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.

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11.3

Specific Gravity Data

The specific gravity database for the main Rainy River deposit (excluding the Intrepid Zone) 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

The specific gravity database for the Intrepid Zone includes 661 measurements completed by ALS by pycnometry on pulverized core samples selected as representative of each modelled geological domain (Table 11-2).

Table 11-2: Summary Statistics of Specific Gravity Data for the Intrepid Zone


Zone	Minimum		Maximum		Mean		Standard Deviation		Sample Variance		Count
Low grade		2.63		3.17		2.84		0.08		0.01		72
Medium grade		2.62		2.95		2.82		0.07		0.01		85
High grade		2.64		3.03		2.82		0.09		0.01		94
All ore		2.62		3.17		2.83		0.08		0.01		251
Waste		2.55		4.48		2.82		0.12		0.01		410

There are insufficient specific gravity data to model the spatial variance of specific gravity across the deposit area. Accordingly, SRK assigned an average specific gravity value of 2.82 to all resource domains to convert volumes into tonnages.

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11.4

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 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 on the main Rainy River deposit and between January 2012 and June 2013 on the Intrepid Zone.

Between December 2011 and June 2013, 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.

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Eight (8) commercial certified reference material control samples were used during this period, sourced from CDN Resource Laboratories Ltd. (“CDN”), Canada.

11.5

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 Chapter 12.

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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), these re-sampled core assays compare well with original assay results reported by Nuinsco.

12.2

Verifications by Rainy River

As outlined in Chapter 11, 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 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, CAP and Intrepid 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 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 relating to the main Rainy deposit, as well as quality control samples for the period of January 2012 to June 2013 relating to the Intrepid Zone.

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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 for the main Rainy River deposit 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 for the Main Rainy River Deposit

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

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.

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Paired data for field duplicates show that gold grades can be reasonably reproduced by ALS Minerals. 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 Minerals 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.

The external analytical quality control data produced by Rainy River for the Intrepid Zone are summarized in Table 12-2 and presented in graphical format in Appendix D. The external quality control data produced on this project represents approximately 4 percent of the total number of samples.

Table 12-2: Summary of Analytical Quality Control Data Produced

between January 2012 to June 2013 for the Intrepid Zone


	Total		(%)			Comment
Sample count		41,020
Field blanks		235		0.57	%	Coarse marble
Certified standards		1,162		2.83	%
26		44				CDN (0.372 ± 0.048 gpt gold)
P3B		248				CDN (0.409 ± 0.042 gpt gold)
P4A		3				CDN (0.438 ± 0.032 gpt gold)
IJ		203				CDN (0.946 ± 0.102 gpt gold)
IH		4				CDN (0.972 ± 0.108 gpt gold)
1L		48				CDN (1.16 ± 0.10 gpt gold)
1P5K		11				CDN (1.44 ± 0.13 gpt gold)
1P5D		4				CDN (1.47 ± 0.15 gpt gold)
1P5E		281				CDN (1.52 ± 0.11 gpt gold)
5H		51				CDN (3.88 ± 0.28 gpt gold, 50.4 ± 2.7 gpt silver)
5G		2				CDN (4.77 ± 0.40 gpt gold, 101.8 ± 72.5 gpt silver)
5J		263				CDN (4.96 ± 0.42 gpt gold, 7.0 ± 4.8 gpt silver)
Field duplicates		224		0.55	%
Total QC samples		1,621		3.95	%
Check assays
ALS Minerals vs. Actlabs		317		0.77	%

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In the opinion of SRK, the results of the analytical quality control data produced from December 2011 to June 2013 are sufficiently reliable to support resource estimation.

12.3.3

Verification of Electronic Data

Rainy River made a Gemcom database available to SRK containing all electronic data accumulated on the Rainy River Project. SRK conducted a series of routine verifications to 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 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 published in 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 that would be low in sulphide content 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

The results presented in Sections 13.2 to 13.14 were used as the basis for the Feasibility Update Study. The testwork was completed from October 2011 until November 2012 on the main pit and from November 2012 to November 2013 on the Intrepid Zone. The objective of the Intrepid Zone testwork program was to confirm that this material could be treated using the proposed flowsheet for the main pit and would not have a significant impact on the design of the plant when blended at low tonnages.

Composite Selection Three (3) composites were used for determining the flowsheet and the parameters for the variability testwork:

•

ODM Master;

•

Initial Pit;

•

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, RLOM Composites and Overall Pit


Zone	Initial Pit (%)		RLOM (%)		Overall Pit (%)
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

A master composite for the Intrepid Zone was also generated using the variability samples.

13.2.1

Variability Sample Selection

In order to provide a high level of confidence, 162 comminution and 208 leaching samples were selected for variability testwork for the main pit and another 30 comminution and leaching samples were selected for the Intrepid Zone. 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 in the main pit are presented in Figure 13-1 and Figure 13-2.

LOGO

Figure 13-1: Sample Locations for Comminution Variability Testwork

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LOGO

Figure 13-2: Sample Locations for Metallurgical Variability Testwork

The sample selection provides a good representation of the entire proposed pit.

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 13-3, where the August 2012 pit outline is in purple and the December 2011 pit is in grey.

LOGO

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

Variability Testwork (Cross-Section View, Looking West)

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13.2.2

Sample Characterization

A large number of samples from each of the zones were selected and assayed for the variability testwork campaign, as described in Section 13.2.1. Three (3) master composites were also assayed: Initial Pit, Remaining Life of Mine and Intrepid Zone master composites.

The head grades and major impurities for the variability samples and Master Composite are presented in Table 13-3.

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Table 13-3: Testwork Sample Head Grades

Sample

# of
Samples
-

Au
(Screen Met.)

g/t

(Direct)

g/t

Cu
%

S
%

Zn
%

Variability Samples¹

ODM

117

1.04 +/- 1.59

4.04 +/- 4.18

0.010 +/- 0.012

2.07 +/- 0.88

0.13 +/- 0.15

2.74 +/- 0.87

Z433

1.12 +/- 1.21

2.03 +/- 1.25

0.041 +/- 0.036

2.22 +/- 1.70

0.06 +/- 0.09

4.20 +/- 1.82

0.51 +/- 0.32

1.00 +/- 0.71

0.015 +/- 0.008

2.15 +/- 0.94

0.06 +/- 0.03

3.23 +/- 0.63

0.79 +/- 0.84

1.99 +/- 0.92

0.019 +/- 0.020

2.25 +/- 0.90

0.07 +/- 0.11

3.54 +/- 0.79

CAP

0.72 +/- 0.82

3.65 +/- 0.71

0.017 +/- 0.015

3.70 +/- 1.97

0.07 +/- 0.08

9.39 +/- 3.44

Intrepid Zone

1.64 +/- 1.79

14.9 +/- 17.4

0.009 +/- 0.009

2.27 +/- 0.42

0.11 +/- 0.16

2.37 +/- 0.40

Master Composites

Initial Pit

—

0.90

2.57

0.016

2.05

0.15

3.13

Remaining Life of Mine

—

0.71

2.86

0.010

2.54

0.07

4.05

Intrepid Zone Master

—

1.45

13.8

0.009

2.19

0.10

2.34

^1.	Numbers shown are averages +/- one standard deviation (s).

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The samples from the CAP Zone have significantly higher levels of sulphur than any of the other zones. The iron levels in the CAP Zone are also considerably higher than any other zone. The Intrepid Zone has much higher silver levels than the other zones; however the copper, sulphur, zinc and iron levels seem to correspond well to the samples from the main pit (excluding the CAP Zone samples).

It can be also seen that the Master Composite for the Intrepid Zone corresponds well in terms of the elemental averages of the variability samples, and Initial Pit and Remaining Life of Mine composites correspond well to the averages of the zones of the main pit.

13.3

Mineralogy

Gold deportment studies have been performed on samples from each zone during the 2011 and 2012 testwork campaigns, including: five ODM, two 433, one CAP, one HS and one 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 433 Composites to 9.5% in the CAP composite).

–

Gold associated with pyrite can be recovered and concentrated using froth flotation;

–

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;

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–

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 of 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.

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”);

•

Gravity Tailings Leach: This flowsheet is a whole rock cyanide leach process followed by CIP.

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.

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A gravity circuit is included in the grinding circuit for both flowsheets. The gravity circuit was included based on the moderately high gravity recoveries noted during historical testwork. Testwork for the gravity circuit for the flowsheet selection were 10 kg tests performed on the sample prior to flotation or leaching testwork. These tests were performed using a Knelson concentrator and the concentrate was upgraded with a Mozley separator. The results from the gravity separation testwork are presented along with the flotation and leach results.

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 13-4.

Table 13-4: Rougher Flotation Results


						Assays						% Distribution
	Combined											Au				Ag
Test No.	Products K₈₀, µm		Product	Mass %		Au (g/t)		Ag (g/t)		S⁼ (%)		Flot (unit)		Grav+ Flot		Flot (unit)		Grav+ Flot		S⁼
1			Ro Conc 1-5		14.6		4.33		18.3		12.9		91.9		94.3		77.6		78.7		97.4
		163	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			Ro Conc 1-5		14.6		4.03		19.5		13.0		92.0		94.3		78.7		79.7		97.8
		101	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			Ro Conc 1-5		20.6		2.70		13.7		9.3		90.9		93.6		78.1		79.1		97.2
		79	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			Ro Conc 1-5		19.6		3.38		14.6		10.2		92.7		94.8		79.9		80.8		98.0
		76	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%.

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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 13-5.

For the first test the concentrate was reground. In the second test, the concentrate was floated without regrinding.

Table 13-5: Cleaner Flotation Results


				Assays						% Distribution
	Product	Mass %		Au (g/t)		Ag (g/t)		S⁼ (%)		Au				Ag
										Au				Ag
Feed Size P₈₀, µm										Flot		Grav+ Flot		Flot		Grav+ Flot		S⁼
Feed Size P₈₀, µm										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 13-6 and Table 13-7.

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


	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 13-7: Flotation Concentrate Leaching Results (Silver Assays)


	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 1.2 to 4.9 kg_NaCN/t and 0.8 to 1.8 kg_CaO/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 it 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 13-8.

Table 13-8: Flotation Tailings Leaching Results


	Feed	Feed		Reag.
	(Tails	Size		Consumption				% Au Recovery
	from	P₈₀,		kg/t of CN Feed				CN Leach Time (h)										Residue		Head Au
CN Test No.	Test)	µm		NaCN		CaO		6		24		36		48		Rec.¹		Au (g/t)		(g/t)
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 indicate additional overall recovery.

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 13-8 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.

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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 (S.G. of 3.06 t/m³) 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 13-9.

Table 13-9: IsaMill Testwork Results


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

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 13-4.

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LOGO

Figure 13-4: Specific Energy vs. Particle Size for IsaMill Testing

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 13-10.

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Table 13-10: 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 13-5.

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LOGO

Figure 13-5: Specific Energy vs. Particle Size for SMD Testing

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 that 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 13-11 and Table 13-12, respectively. Any results presented for grind sizes with more than one (1) test are averaged values.

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Table 13-11: Gravity Tailings Leach Results (Gold)


			Reagent				Au Recovery (%)												Au Assays
Number of tests	P₈₀ (µm)		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 13-12: Gravity Tailings Leach Results (Silver)


			Reagent																Ag Assays
Number of tests			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
1*		69		0.09		0.37		64.9		75.4		74.2		78.8		6.7		80.2		0.70		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

*	One test result not presented as silver residue levels were much lower than any other test.

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% (1% higher than historical testwork) 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. The gravity circuit was included due to relatively high gravity recovery (25-37%) noted during the testwork. 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 13-13.

Table 13-13: 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^th Percentile		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^thpercentiles, 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 on the main pit, along with an additional two samples from the Intrepid Zone. 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. Using this method, 162 total samples (from the main pit) were selected for SMC testing as described in section 13-3 in addition to the 13 samples that the JK DWT tests were performed on, for a total of 175 SMC tests.

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The JK DWT results consist of A and b factors (often referenced as A x b) that indicate the resistance to impact breakage and a t_a value, which indicates the resistance to abrasion. A lower A x b value indicates a higher resistance to impact breakage while a higher t_a value indicates a material that is more resistant to abrasion breakage. 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 13-14.

Table 13-14: JK Drop Weight and Corresponding SMC Results


	JK DWT										SMC
Zone	A		B		A x b		t_a		r (g/cm³)		A		b		A x b		A x b % 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
Intrepid		65.9		1.36		89.6		1.04		2.63		64.9		1.60		104		16.1
Intrepid		100		0.23		23.0		0.28		2.72		83.4		0.60		38.0		65.2
														Average (Main pit)				5.1
										Average (including Intrepid Zone)								9.8

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It can be seen from Table 13-14 that the calibrated SMC test results are slightly higher than the JK DWT results for the same sample, indicating that for a given sample the SMC results will indicate slightly less resistant to impact breakage than the JK DWT. 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 multiple SMC tests.

The differences on the Intrepid Zone samples between the JK DWT and SMC tests were more significant than the other zones. One of the samples tested had a much higher A x b value than any other samples tested from the main pit. This is an indication that the ore hardness variability in the Intrepid Zone may be more significant than the zones from the main pit. However, this is not expected to have a significant impact on operations given the high blending ratio that will be used when introducing the Intrepid Zone material to the mill.

As can be seen in Figure 13-6, the JK DWT tests (indicated as diamonds) fall onto the curve generated by the multiple SMC tests, as expected.

LOGO

Figure 13-6: SMC Data Distribution with JK DWT Calibration Points

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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 13-15. It should be noted that a lower A x b or a higher M_ianumber is indicative of a harder material.

Table 13-15: SAG Mill Comminution SMC and M_ia Values


Zone	ODM		Z-433		HS		NZ		CAP		Waste
A x b
Number of Tests		95		19		12		21		26		2
Average		32.9		23.7		22.0		28.3		23.2		21.6
Minimum		62.6		38.6		24.9		63.3		34.7		22.0
Median		32.4		22.7		22.1		26.0		22.3		21.6
80^th Percentile		26.6		20.7		20.8		21.8		20.3		21.3
Maximum		20.7		19.0		19.0		20.0		18.0		21.1
M_ia (kWh/t)
Average		23.6		30.0		31.5		27.0		30.3		31.9
Minimum		13.8		19.9		28.5		13.5		21.6		31.4
Median		23.2		30.4		31.1		27.4		30.6		31.9
80^th Percentile		27.0		32.5		33.0		31.4		33.2		32.1
Maximum		32.8		35.6		35.2		34.6		37.4		32.3

Based on reference and industrial data, 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 80^th 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 t_a 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 13-16.

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Table 13-16: 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_SAGto 1.7mm (kWh/t)		SAG Dis. Bond BWI (kWh/t)		Calc W_BMto P₈₀(kWh/t)		Total W_Tto 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
Minimum			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
Maximum			7499		1770		1.67		2.97		13.3		14.8		13.5		26.8
Design Data (80th Percentile)									2.81		12.2		14.6		13.3		25.4

The W_SAG values provided by SAG Design ranged from 9.5 to 13.3 kWh/t with an 80^th percentile value 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 analysis of the SAGDesign method compared to the JKDWT method.

A comparison of the SAGDesign and JK DWT test program results is presented in Table 13-17.

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Table 13-17: 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.

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 on material from the main pit. 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 13-7.

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LOGO

Figure 13-7: 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 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 13-18.

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Table 13-18: Bond and ModBond Results


Description	Bond Work Index (200 mesh) kWh/t
Zone	ODM		Z-433		HS		NZ		CAP		Intrepid
Bond Work Index, 200 mesh (kWh/t)
Number of Tests		5		4		2		2		3		8
Average		13.6		15.6		16.2		13.0		15.2		16.7
Minimum		12.6		15.2		16.1		12.1		14.8		13.2
Median		13.8		15.7		16.2		13.0		14.9		15.6
80^th Percentile		14.2		15.9		16.2		13.5		15.6		19.0
Maximum		15.0		15.9		16.3		13.8		16.1		21.5
ModBond Work Index, 200 mesh (kWh/t)
Number of Tests		89		17		10		20		24		30
Average		13.8		15.1		14.9		14.1		14.7		15.1
Minimum		11.6		12.9		14.1		11.1		13.0		13.4
Median		13.8		15.3		15.0		14.2		14.8		15.1
80^th Percentile		14.7		15.4		15.2		15.0		15.5		15.7
Maximum		16.0		15.8		15.5		16.2		15.8		17.3

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 results include results from testwork for the Intrepid Zone that was performed after completion of the testwork program for the main zone. It can be seen that for the Intrepid Zone, at the 80^th percentile, the BWi and ModBWi are considerably different, with values of 19.0 and 15.7 kWh/t, respectively. This is due to two (2) samples that had considerably higher BWi values. When ignoring these two (2) results, the 80^th percentile of the BWi tests is 15.7 kWh/t –identical to the ModBWi results. Overall the Intrepid Zone has slightly higher BWi and ModBWi values, indicating that the zone is harder than the zones from the main pit.

The distribution of the ModBond work indices can be seen in Figure 13-8.

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LOGO

Figure 13-8: 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 15.7 kWh/t. A weighted average value of 15.0 kWh/t was used for design.

13.5.6

ModBond and A x b

Figure 13-9 shows an interesting behaviour between A x b and ModBond for the main pit. 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 13-9: ModBond Wi vs. A x b

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 13-19.

Table 13-19: 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^th Percentile (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 when compared to the SGS database.

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13.5.8

Grinding Circuit Design

Several different design methods were used to size the SAG and ball mill circuit. The 80^thpercentile 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 and consultants were used to estimate the size of the grinding circuit:

•

Morrell’s Equation: This method uses the M_ia and 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_a results 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) that 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.

•

Orway 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 P₈₀ of 75 µm;

•

A x b value of 24.2, t_a value of 0.35;

•

BWi value of 15.0 kWh/t; and

•

M_ia value of 29.3 kWh/t.

The results from the simulations are presented in Table 13-20.

Table 13-20: SAG and Ball Mill Simulation Results¹


			Morrell’s Equation		JKSimMet + Bond’s Equation		JK SimMet + Phantom Cy		SAGDesign		OMC
Method	Units		80^th Percentile		80^th Percentile		80^th Percentile		79^th Percentile²		80^th Percentile
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.	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 that particle sizes have not been adjusted for 21,000 tpd.

^2.	The 79th percentile used for the SAGDesign simulations was based on seven (7) samples only.

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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 JKSimMet 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, 15 MW dual pinion drives were selected for both the SAG mill and ball mill for a mill fresh 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. Based on the results, a 36’ x 20’ (18.25’ EGL) SAG mill and a 26’ x 40.5’ (40’ EGL) ball mill were selected based on 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.

Subsequent simulations performed at 21,000 tpd (951 tph) indicated that the T₈₀ of the SAG Mill circuit would be 2,800 µm rather than 2,400 µm. This is not expected to have a significant impact on the design of the plant.

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 433 Zones. The test results are presented in Table 13-21.

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Table 13-21: Gravity Recoverable Gold Results


	Grind Size		Mass				Assay Au		Au grams		Dist’n
Sample	P₈₀ (µm)	Product	grams		%		(g/t)		(g)		%
	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
ODM Master	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
	Final Conc			254		1.3		39.7		10,065		51.2
	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
Z433	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
	Final 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

Gravity Variability 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 excluding the Intrepid Zone. The gravity gold recovery from the CAP Zone was considerably lower with an average recovery of 9%. The Intrepid Zone also had considerably lower recoveries, averaging 16% gold recovery by gravity.

The gold and silver recoveries as a function of head grade are presented in Figure 13-10 and Figure 13-11.

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LOGO

Figure 13-10: Gold Gravity Recovery vs. Head Grade

LOGO

Figure 13-11: Silver Gravity Recovery vs. Head Grade

It can be seen that the gravity recovery of gold is sensitive to the head grade, with higher grades giving higher recoveries. The same trend was noted for the Non-CAP, CAP and Intrepid Zones; however, the CAP and Intrepid Zone gold recovery were 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 recoveries for the CAP and Intrepid Zones were lower than the non-CAP Zones, analogous to the gold gravity recovery. The average silver gravity recovery for the CAP Zone was roughly 3%; the Intrepid Zone averaged 5% while the non-CAP Zones silver gravity recovery was approximately 10%.

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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 13-12.

LOGO

Figure 13-12: 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 measured for 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, 18, 24 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)				Reagent Consumptions (kg/t)														Ag Assays
																									Head
	Number		P₈₀																		Grav +		Residue		Grade
Comp Name	of Tests		(µm)		NaCN		CaO		12h		18h		24h		30h		36h		Grav		CN		(g/t)		(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
																									Head
	Number		P₈₀																		Grav +		Residu		Grade
Comp Name	of Tests		(µm)		NaCN		CaO		12h		18h		24h		30h		36h		Grav		CN		e (g/t)		(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 13-13. The dotted lines represent the sensitivity of the assay technique (0.02 g/t).

LOGO

Figure 13-13: 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 13-13. 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 13-14.

LOGO

Figure 13-14: 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.

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 described in Section 13.8.7.

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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. A consistent trend between cyanide concentration and gold/silver recovery was not observed.

The results are plotted in Figure 13-15.

LOGO

Figure 13-15: Gold Recovery vs. Time at Different NaCN Concentrations

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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.

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 preconditioning prior to leaching.

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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.

Table 13-26: O₂ vs. Air and Lead Nitrate Addition Testwork Results


						Reagent Consumptions (kg/t)				Au Recovery (%)												Au Assays
Comp Name	Aeration		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		0.97
Initial Pit		O₂	Y		59		0.16		0.50		—		—		—		88		29.0		91.5		0.08
		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
		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		0.89
		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.

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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.

Based on these results, it was decided that the variability tests would be done without lead nitrate.

13.8.6

Intrepid Zone Leaching Kinetics

Given the large amount of silver in the Intrepid Zone, the leaching kinetics for both gold and silver were investigated. To evaluate the silver leaching kinetics, the following conditions were considered:

•

Leach time of 96 hours with a sub-sample at 30, 36, 48 and 72 hours;

•

Target grind size (P₈₀) of 75 µm;

•

Cyanide concentration of 0.5 g/L NaCN;

•

30-minute pre-conditioning;

•

pH of 10.5-11.

The first two (2) tests (as well as a repeat of the first test) were performed without gravity concentration, while the subsequent two (2) tests included gravity separation.

The cyanide leaching kinetics results are presented in Table 13-27 and Table 13-28.

Table 13-27: Intrepid Zone Leaching Kinetics Tests (Gold)


					Reagent Consumptions (kg/t)						Au Recovery (%)												Au Assays
Test	P₈₀ (µm)		NaCN Conc. (g/L)		NaCN		CaO		30 h		36 h		48 h		72 h		96 h		Grav		Total		Residue (g/t)		Head Grade (g/t)
1		78		0.5		0.19		0.36		88		89		90		88		89.3		n/a		89.3		0.17
1R		80		0.5		0.18		0.31		91		90		93		95		95.5		n/a		95.5		0.10
2		74		1.0		0.31		0.32		92		93		93		92		92.2		n/a		92.2		0.13		1.45
3		74		0.5		0.20		0.37		91		92		93		92		90.9		16.9		92.4		0.13
4		72		1.0		0.30		0.33		93		93		93		92		92.1		16.9		93.4		0.12

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Table 13-28: Intrepid Zone Leaching Kinetics Tests (Silver)


					Reagent Consumptions (kg/t)				Ag Recovery (%)														Ag Assays
Test	P₈₀ (µm)		NaCN Conc. (g/L)		NaCN		CaO		30 h		36 h		48 h		72 h		96 h		Grav		Total		Residue (g/t)		Head Grade (g/t)
1		78		0.5		0.19		0.36		56		58		62		68		70.1		n/a		70.1		4.3
1R		80		0.5		0.18		0.31		52		55		60		67		69.7		n/a		69.7		4.6
2		74		1.0		0.31		0.32		58		61		64		70		72.5		n/a		72.5		4.1		13.8
3		74		0.5		0.20		0.37		50		52		56		64		67.5		3.8		68.7		4.5
4		72		1.0		0.30		0.33		60		61		65		73		74.5		3.8		75.5		3.6

The average leaching extraction for gold and silver was plotted as a function of time to analyse the reaction kinetics. The standard deviations for each point were also shown. This data includes results with and without gravity separation.

This data can be seen in Figure 13-16.

LOGO

Figure 13-16: Gold and Silver Cyanide Leaching Kinetics (Intrepid Zone)

The silver kinetics for the Intrepid Zone are quite slow, with increases in silver extraction noticeable up to 96 hours, and possibly longer. The gold extraction kinetics are considerably faster, with little to no improvement noted from 30 hours to 96 hours. Based on this information, it was determined that material from the Intrepid Zone would be treated using the same leaching conditions as the main pit.

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13.8.7

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 for the main pit. An additional 30 samples from the Intrepid Zone were also tested.

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 13-29 and Table 13-30.

Table 13-29: Gold Leaching Variability Testwork Average Results


					Reag. Consumption				% Au Recovery
	# of		P₈₀		(kg/t)				CN (Unit)								Overall		Residue		Direct Au
Zone	Tests		(µm)¹		NaCN		CaO		30 h		36 h		Grav		Overall²		Recalculated³		Au (g/t)		Head (g/t)¹
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
Intrepid		30		75 +/- 6		0.10		0.31		86.1		86.9		16.1		88.0		92.0		0.13		1.57 +/- 1.79
*Non-CAP Subtotal*		*235*		*89 +/- 36*		*0.07*		*0.37*		*80.5*		*81.3*		*25.7*		*85.7*		*90.8*		*0.10*		*1.17 +/- 1.65*
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		275		89 +/- 38		0.08		0.41		79.1		79.9		23.4		84.0		89.8		0.11		1.09 +/- 1.56

^1.	Values shown are averages +/- the standard deviation (s) of the results.

^2.	The overall recovery numbers are artificially low as low grade samples have same weighting as high grade samples.

^3.	Gold recoveries recalculated using average direct head grade and average residue values.

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Table 13-30: Silver Leaching Variability Testwork Average Results


					Reag. Consumption				% Ag Recovery
	# of		P₈₀		(kg/t)				CN (Unit)								Overall		Residue		Direct Ag
Zone	Tests		(µm)¹		NaCN		CaO		30 h		36 h		Grav		Overall²		Recalculated³		Ag (g/t)		Head (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
Intrepid		30		75 +/- 6		0.10		0.31		58.5		60.2		5.3		60.5		55.5		6.6		14.9 +/-17.4
*Non-CAP Subtotal*		*235*		*89 +/- 36*		*0.07*		*0.37*		*55.8*		*57.2*		*9.5*		*60.9*		*61.4*		*1.73*		*4.48 +/-9.32*
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		275		89 +/- 38		0.08		0.41		57.0		58.3		8.6		61.9		61.9		1.60		4.21 +/-8.65

^1.	Values shown are averages +/- the standard deviation (s) of the results.

^2.	The overall recovery numbers are artificially low due as low grade samples have same weighting as high grade samples.

^3.	Silver recovery recalculated using average direct head grade and average residue values

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 Non-CAP zones (including the Intrepid Zone) 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 89 µm was considerably coarser than the target P₈₀of 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.

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The filtered gold and silver residues are plotted as a function of head grade in Figure 13-17 and Figure 13-18, respectively. The Non-CAP curves do not include data from the Intrepid Zone as the numbers of samples tested per tonne of material is considerably higher than the other zones and might influence the recovery curves. Given that Intrepid Zone will be blended at low tonnage rates and appears to behave similarly to Non-CAP material, it is not anticipated that the Intrepid Zone samples will have any impact on the Non-CAP recovery curve.

LOGO

Figure 13-17: Gold Residue vs. Head Grade (Variability Tests)

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LOGO

Figure 13-18: Silver Residue vs. Head Grade (Variability Tests)

Figure 13-17 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.

13.8.8

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

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•

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.

13.8.9

Mercury Assays

Mercury assays were performed on two composite samples, measuring mercury levels in the feed, residue, loaded carbon and barren solution streams after undergoing leaching and gold adsorption (CIP). The objective of the testwork was to determine if any mercury leached into solution and adsorbed onto the carbon. All assays were below detection level except for one carbon reading, which had an assay of 0.06 g/t mercury.

13.9

Cyanide Destruction Testwork

The SO₂/air cyanide destruction process was investigated on three (3) composites: Initial Pit, Remaining-Life-of-Mine and Intrepid Zone. The Intrepid Zone sample was tested after completion of the main pit testwork. The first series of tests on the Intrepid Zone sample yielded high residual cyanide levels, however a repeat of the test showed results in line with those from the main pit samples. One (1) large bulk cyanide destruction and three (3) continuous tests were conducted for each composite.

The cyanide testwork results are presented in Table 13-31.

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Table 13-31: 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
	Feed		—		—		10.7		151		77.4		—		20.0		2.22		—		—		—
Intrepid Zone	Batch
	CND 2		50		150		8.6		—		—		0.26		—		—		11.9		7.68		0.13
	Continuous
	CND 2-1		50		58		8.5		0.13		<0.1		4.1		18.0		0.20		4.64		2.36		0.13
	CND 2-1		50		116		8.6		0.11		<0.1		0.94		7.30		0.20
	CND 2-2		50		58		8.5		<0.1		<0.1		0.45		5.10		0.30
	CND 2-2		50		116		8.5		<0.1		<0.1		<0.1		1.10		0.20		5.69		3.64		0.12

^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 for the main pit, respectively. The reagent consumptions for the Intrepid Zone sample were comparable to these values. These consumptions are considered to be 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.

The Initial Pit, Remaining-Life-of-Mine and Intrepid Zone master composites were used for the CIP modeling testwork.

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The isotherms from the testwork are presented in Figure 13-19.

LOGO

Figure 13-19: 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 that 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, 79 and 90 for the Initial Pit, Remaining-Life-of-Mine and Intrepid Zone 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 0.007 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 three (3) 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 13-32.

Table 13-32: Flocculant Description


		Molecular Weight		Charge Density
Flocculant	Charge	(10⁶ Dalton)		(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 13-33.

Table 13-33: Sedimentation Testwork Results


	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		951		951		951		951	951	951
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		45		39		39		39	46	46

The results indicated that the recommended thickener diameter is between 39 and 46 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 45 m diameter pre-leach and pre-detox thickeners be used.

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.

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Intrepid Zone Testwork

Static and dynamic settling tests were also performed at SGS on the Intrepid Zone material. The results indicate that the thickening of the material from the Intrepid Zone is relatively slow. This was noted by the solids loading rates of 0.42 and 0.61 t/h/m². The flocculant addition rate (25 g/t average in the dynamic tests and 20 g/t in the static tests) is in-line with industrial practices. Good overflow clarity was noted in both the static and dynamic tests. The Intrepid Zone material is not expected to have an impact on design.

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 the 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 13-20 and Figure 13-21.

LOGO

Figure 13-20: Initial Pit Composite Yield Stress vs. Solids Density

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LOGO

Figure 13-21: 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 61% and 60%, respectively.

13.13

Linear Screen Sizing Testwork

Testwork was performed at a supplier laboratory to determine the sizing of the trash and safety linear screens. This was done by determining the flux rate (m³/m² /h) at different feed percent solids. The results are presented in Table 13-34.

Table 13-34: Linear Screen Sizing Testwork Results


Description	Units	Trash Screen					Safety Screen
Feed Volume	m³/h		2,777					1,296
% Solids	% w/w		28	28	28	28		50
Screen Aperture	µm		600	600	700	700		600
Draining Rate (Flux Rate)	m³/m ²/h		88	88	98	98		51
Required Screen Area	m²		31.6	31.6	28.3	28.3		25.3
Design Safety Factor	%		20	20	20	15		20
Required Screen Area	m²		37.9	37.9	34.0	32.6		30.4
Number of Screens			1	2	1	1		1
Screen Size	m²		40	20	40	40		32

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The testwork indicated a flux rate to the trash screen of 88 and 98 m³/m²/h for screen apertures of 600 and 700 µm, respectively. The safety screen indicated a flux rate of 51 m³/m²/h for a 600 µm screen aperture. Based on these results, two (2) 20 m² trash screens and one (1) 32 m²safety screen were selected with screen apertures of 600 µm.

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 13-35 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.

Table 13-35: Summary of Geochemical Environmental Testing


Analytical Procedure	Mine Rock		Overburden		Tailings
Static Testing
Acid-base Accounting¹		362		3		6
Net Acid Generation test		176		3		6
Metals Content (aqua-regia ICP)		362		3		6
Short-term Leach testing (SFE²)		74		2		6
Mineralogy (X-Ray Diffraction)		39		0		6
Kinetic Testing
Humidity Cell		20		0		6
Field Cell		7		0		0

^1.	Methodology: Sobek with siderite correction, modified Sobek, total inorganic carbon and sulphur speciation.

^2.	SFE = 24 hour shake flask extraction at 3:1 deionized water to solid.

Geochemical studies and bock modeling results indicate that approximately one-half (51%, or 163Mt) of the mine rock waste is potentially acid generating. 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. However, testing on tailings samples has indicated a potential for release of cadmium and zinc. Laboratory humidity cell tests on mine rock and tailings and in-field barrel tests on mine rock are currently being undertaken to evaluate the long-term acid generation potential and metal leaching characteristics of these materials.

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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 13.8.7 and the proposed whole rock leach flowsheet, discussed in Chapter 17. The recovery curves were developed using the following format:

LOGO

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, HS and Intrepid). These curves are presented in Table 13-36.

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Table 13-36: Residue and Gravity Recovery Curves

LOGO

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.

Using the inputs presented above, the gold and silver recovery curves were simplified and are presented in Table 13-37.

Table 13-37: Simplified Gold and Silver Recovery Curves

LOGO

The expected gold and silver recoveries as a function of head grade (gold or silver, respectively) are presented in Table 13-38.

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Table 13-38: 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 Update Study. Based on a trade-off study, it was determined that the whole rock leaching option with gravity separation was the most economical alternative and was therefore used as the basis for the Feasibility Update 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 A x b value was modified from 34.0 to 24.2, resulting in an

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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.

Testwork was performed on samples from the Intrepid Zone after completion of the testwork on the main pit. The objective of this testwork program was to confirm that the Intrepid Zone could be treated using the proposed flowsheet and would not have a significant impact on the design of the plant when blended at low tonnages. The Non-CAP Zone recovery curve was determined during the main pit testwork program without the inclusion of the Intrepid Zone material. The test work performed on the Intrepid Zone indicates that the material should not impact design if blended at low tonnage rates. Given the high silver grade of this material, it is not recommended to be blended at high tonnage due to downstream effects on CIP and elution.

A gravity circuit was included in the plant design due to the good GRG results (51.2 and 59.3) along with the moderate gravity recoveries noted for the Non-CAP zones during the variability testwork (average of 25.7%).

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. When calculating gold recoveries, the material has been divided into Non-CAP zones (ODM, Z-433, NZ, HS and Intrepid) and CAP Zone.

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 - 8.5, resulting in an elevated recovery for those years.

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Average gold and silver recoveries are based on the mine plan and shown in Table 13-39.

Table 13-39: Average Yearly Gold and Silver Recoveries


	Gold				Silver
Years	Au Head Grade (g/t)		Au Recovery (%)		Ag Head Grade (g/t)		Ag Recovery (%)
1		1.44		91.9		2.13		65.5
2		1.48		92.1		2.00		65.7
3		1.45		92.0		3.80		63.1
4		1.43		91.9		3.56		63.5
5		1.38		91.7		4.92		61.5
6		1.46		92.0		4.27		62.4
7		1.50		92.1		2.74		64.6
8		1.57		92.3		2.15		65.5
9		1.21		91.0		2.04		65.6
10		0.68		87.3		1.64		66.2
11		0.60		86.2		2.59		64.8
12		0.40		82.3		2.81		64.5
13		0.33		80.2		2.11		65.5
14		0.55		74.2		2.33		64.8
LOM		1.12		90.6		2.80		64.1

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14.

MINERAL RESOURCE ESTIMATION

14.1

Introduction

The resource estimation work was completed in Toronto by Dorota El-Rassi, P.Eng (PEO #100012348), assisted by Mr. Bob Kusins, P.Geo. (APGO #0196) and under the supervision of Glen Cole, P.Geo. (APGO #1416), all “independent qualified persons”, as this term is defined in National Instrument 43-101.

The Mineral Resource Statement presented herein represents the ninth (9^th) mineral resource evaluation prepared for the Rainy River Project since 2003 that includes, for the first time, the Intrepid Zone. The Intrepid Zone is part of the Rainy River Project, located approximately one kilometer east of the main Rainy River deposit.

The Rainy River 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 and silver content only, although locally significant silver is also present. The consolidated Mineral Resource Statement for the Rainy River 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 ninth (9^th) Mineral Resource Statement for the Rainy River Project.

In the opinion of SRK, the resource evaluation reported herein is a reasonable representation of the mineral resources found in the Rainy River 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 entire Rainy River 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 strategies 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.

14.1

Resource Database

Exploration data used to evaluate the mineral resources for the Rainy River Project were provided by Rainy River as a set of Microsoft Excel files containing drilling information (drill collars, surveys, assays and lithological logging) for 1,665 core boreholes (742,424 m) drilled by Rainy River and Nuinsco, the previous Project operator. The database includes 230 boreholes (79,575 m) drilled within the Intrepid Zone during the period from 1996 to August 16^th, 2013, 237 boreholes (95,760 m) drilled in 2012, 388 boreholes (188,588 m) drilled by Rainy River in 2011 and early 2012, 165 boreholes (84,134 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 topographic 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 for the main Rainy River deposit between December 2011 and June 2012. For the Intrepid Zone, analytical quality control data was considered for the period from December 2011 to June 2013. 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 by BBA with MineSight software.

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14.3

Solid Body Modelling

14.3.1

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 2012 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. With the discovery of the Intrepid Zone, Rainy River, with SRK’s assistance, constructed a separate 3D domain model expanding the Rainy River Project to the east (Figure 14-2).

The Rainy River Project was subdivided into 13 separate resource domains: six (6) main Zones (ODM/17, 34, 433, HS, CAP, Western), the Intrepid Zone, 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 14-3. The ODM/17, 433 and Intrepid Zones were subdivided into three (3) grade subdomains (Low, Medium and High grade).

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Table 14-1: Rock Codes in the Rainy River 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
Intrepid	Low Grade (between 0.3 and 0.8 g/t gold)		100
	Medium Grade (between 0.8 and 2.0 g/t gold)		200
	High Grade (above 2.0 g/t gold)		300
Silver			901 to 904

LOGO

Figure 14-1: Location of New Boreholes Drilled on the Main Rainy River Deposit

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LOGO

Figure 14-2: Location of Drill Collars and High Grade Subdomain for the Intrepid Zone

LOGO

Figure 14-3: Isometric View of the Rainy River Mineralization Wireframes Modelled by SRK with

Borehole Data (View looking towards the West)

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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 14-3) 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.

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.

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14.1.2

The 433, HS and New Zones

The 433 and HS Zones (Figure 14-1 and Figure 14-3) 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:


• 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 14-3 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.

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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.

Western Zone

Gold mineralization appears sporadic in the Western Zone of the Rainy River 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.

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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 contain dendritic native silver inclusions. The Silver Zone domain was outlined by Rainy River using a 19 g/t silver cut-off grade (3.0 m composites; less than 4.0 m waste), on inclined cross-sections perpendicular to the plunge of the silver mineralization.

Intrepid Zone

The Intrepid Zone is located one (1) kilometer east of the ODM17 Zone. Similarly to the main Rainy River Project, the Intrepid Zone contains volcanic hosted gold-rich sulphide mineralization of hydrothermal origin. Rainy River, with SRK’s assistance, constructed a series of three-dimensional wireframes to define the limits of the gold mineralization. The Intrepid Zone was modelled on 17 vertical sections spaced at 25 metres. Three (3) nested grade domains were defined based on the gold and silver content:


• High Grade:		Above 2.0 g/t gold equivalent;

• Medium Grade:		Between 0.8 and 2.0 g/t gold equivalent, and

• Low Grade:		Between 0.3 and 0.8 g/t gold equivalent.

14.4

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 in the Main Rainy River deposit is provided in Figure 14-4. 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.

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LOGO

Figure 14-4: Histogram Distribution of Raw Sample Lengths

14.5

Evaluation of Outliers

SRK constructed cumulative probability curves for the gold and silver composites within each modelled domain (shown in Figure 14-5). 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 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.

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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
Intrepid 100		5		99.7	%		6		60		99.7	%		5
Intrepid 200		20		99.6	%		4		100		99.1	%		11
Intrepid 300		40		99.4	%		5		225		98.9	%		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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LOGO

Figure 14-5: Cumulative Frequency Plot for Gold Composites

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14.6

Statistical Analysis and Variography

14.6.1

Statistical Analysis

The basic statistics for the composite and capped composite data within all the 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
433	Low		300		11,553		0.00		60.19		0.27		1.04		1.08
	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
Intrepid	Low		100		1721		0.01		8.22		0.40		0.60		0.36
	Medium		200		1159		0.01		84.09		1.39		3.40		11.56
	High		300		795		0.02		84.20		3.79		6.38		40.70
600 Series			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
			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
Silver Zones			901		153		0.00		4.04		0.23		0.56		0.31
			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

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Table 14-4: Basic 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
433	Low		300		11,553		0.00		25.00		0.26		0.79		0.63
	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
Intrepid	Low		100		1721		0.01		5.00		0.39		0.53		0.28
	Medium		200		1159		0.01		20.00		1.31		2.07		4.28
	High		300		795		0.02		40.00		3.65		5.13		26.31
600 Series			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
			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
Silver Zones			901		153		0.00		0.80		0.16		0.20		0.04
			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

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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
Intrepid	Low		100		1721		0.05		184.77		6.37		8.64		74.64
	Medium		200		1159		0.05		218.93		15.62		20.08		403.20
	High		300		795		0.05		463.86		30.54		44.30		1965.51
600 Series			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
			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
Silver Zones			901		153		0.00		334.11		59.34		67.10		4502.40
			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

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Table 14-6: Basic 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
Intrepid	Low		100		1721		0.05		60.00		6.24		6.93		48.07
	Medium		200		1159		0.05		100.00		15.16		17.04		290.25
	High		300		795		0.05		225.00		29.50		37.48		1404.67
600 Series			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
			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
Silver Zones			901		153		0.00		250.00		58.17		63.11		3982.66
			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

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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

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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
433	Low		300		11,648		0.01		8.40		1.04		0.85		0.72
	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

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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
433	Low		300		11,648		0.01		7.00		1.04		0.84		0.71
	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

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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
433	Low		300		11,648		0.00		13.37		1.39		1.75		3.05
	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

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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
433	Low		300		11,648		0.00		10.00		1.39		1.73		3.01
	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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14.6.2

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 096° and dips 55° to the south, although local variations do occur. While the majority of the variograms were modelled with two (2) 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 14-6 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.

14-24

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LOGO

Figure 14-6: Examples of Variogram Models for Rainy River Deposit

14-25

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Feasibility Study of the Rainy River Project

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
ODM/17 Zone	Low		100	*		3		0.2		0.2		2		10		15		10		0.3		2		80		60		70
	Medium		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
			112			2		0.3		0.6		2		15		15		7		0.1		1		100		90		50
			113	*		3		0.2		0.5		2		25		0		10		0.1		1		25		90		40
			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
	High		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
			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
433 Zone	Low		300	*		3		0.1		0.4		2		10		30		8		0.25		1		100		45		25
	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
	Low		100East			2		0.2		0.8		1		110		70		3
	Low		100West			2		0.2		0.55		1		50		20		3		0.25		2		60		50		3
Intrepid	Medium		200East			2		0.3		0.4		1		30		40		3		0.3		2		40		80		3
	Medium		200West			2		0.3		0.45		1		20		10		3		0.25		2		80		70		3
	High		300East			2		0.3		0.4		1		60		40		3		0.3		2		80		50		3
	High		300West			2		0.3		0.45		1		40		40		6		0.25		2		50		50		6
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

14-26

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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
Intrepid	Low		100East			1		0.2		0.8		2		110		80		3
	Low		100West			2		0.2		0.55		1		70		135		6		0.25		2		180		135		6
	Medium		200East			2		0.3		0.3		1		30		35		3		0.4		2		110		65		3
	Medium		200West			2		0.3		0.45		1		20		10		3		0.25		2		80		70		3
	High		300East			2		0.3		0.55		1		35		35		3		0.15		2		70		45		3
	High		300West			2		0.3		0.45		1		30		15		3		0.25		2		90		45		8
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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14.7

Block Model and Grade Estimation

14.7.1

Block Model Definition

An unrotated block model was created in Gemcom to cover the entire extent of the Rainy River Project deposit area. Block size was set at 5 m x 5 m x 5 m for the Intrepid Zone and the portion of the main Rainy River deposit amenable to underground mining. The block size for the portion of the Rainy River Project amenable to open pit mining was set at 10 m x 10 m x 10 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 Project Block Model Parameters


Model	Direction	Size (m)		Minimum		Maximum		Number of Cells
	East-West		10		423,700		427,075		338
Open Pit	North-South		10		5,408,750		5,411,125		238
	Vertical		10		(-)1,200		450		165
	East-West		5		423,700		427,075		675
Underground	North-South		5		5,408,750		5,411,125		475
	Vertical		5		(-)1,200		450		330
	East-West		5		427,075		427,675		120
Intrepid	North-South		5		5,409,500		5,409,950		90
	Vertical		5		(-)180		420		120

14.7.2

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 order to test the sensitivity of the grade estimation to the choice of estimation parameters, SRK undertook a series of sensitivity runs varying the interpolation parameters. Results indicate that the models are relatively insensitive to slight variations in the estimation

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parameters. Grade interpolation was completed in two (2) or three (3) successive passes, considering the estimation parameters summarized in Table 14-17 to Table 14-19 for Main Rainy River (amenable to open pit and underground mining) and the Intrepid Zone, respectively. Search neighbourhoods are summarized in Table 14-20 and Table 14-21 for main Rainy River and the Intrepid Zone, respectively.

The search neighbourhoods used for calcium and sulphur interpolation are tabulated in Table 14-22 and Table 14-23, whereas that for Domain 200 (Zone 34) is summarized in Table 14-24. The first estimation pass considered search neighborhoods adjusted to 95% of the variogram sill. The size of the search ellipse was doubled for the second estimation pass.

A third 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 for Part of the Main Rainy River Deposit

Amenable to Open Pit Mining


Interpolation Parameters	1^st Pass (Indicated)		2^nd Pass (Inferred)		3^rd Pass (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		7		5		2
Maximum Number of Composites		12		12		15
Maximum Number of Composites per Borehole		5		3		—
Domains 601 to 605
Interpolation Method		Inverse Distance Power 2		Inverse Distance Power 2		Inverse Distance Power 2
Search Type		Ellipsoidal		Ellipsoidal		Ellipsoidal
Minimum Number of Composites		7		5		2
Maximum Number of Composites		12		12		15
Maximum Number of Composites per Hole		5		3		—
Domain 200 (Pt, Pd, Ni and Cu)
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		7		5		2
Maximum Number of Composites		12		12		15
Maximum Number of Composites per Hole		5		3		—

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Table 14-18: Resource Estimation Parameters for the Part of the Main Rainy River Deposit

Amenable to Underground Mining


Interpolation Parameters	1^st Pass (Indicated)		2^nd Pass (Inferred)		3^rd Pass (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 Composites 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 Composites 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 Composites per Hole		2		—		—

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Table 14-19: Resource Estimation Parameters for the Intrepid Zone


Interpolation Parameters	1^st Pass (Indicated)		2^nd Pass (Inferred)		3^rd Pass (Inferred)
Interpolation Method		Ordinary Kriging		Ordinary Kriging		Ordinary Kriging
Search Type		Octant		Ellipsoidal		Ellipsoidal
Minimum Number of Octants		2		—		—
Maximum Number of Composites per Octant		4		—		—
Minimum Number of Composites		5		3		2
Maximum Number of Composites		10		15		15
Maximum Number of Composites per Borehole		3		2		—

Table 14-20: Search Neighbourhoods Used for Gold and Silver Estimation in Main Rainy River


	Domain		Rotation*						1^st Pass						2^nd Pass						3^rd Pass
			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

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Domain		Rotation*						1^st Pass						2^nd Pass						3^rd Pass
		Rotation*						Search Ranges						Search Ranges						Search Ranges
		Az		Dip		Plunge		X		Y		Z		X		Y		Z		X	Y	Z
	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

Table 14-21: Search Neighbourhoods Used for Gold and Silver Estimation in Intrepid Zone


	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		100West		165		-58		75		60		50		3		160		100		6		180		150		9
		100East		60		58		-60		110		70		3		220		140		6		330		210		9
		200West		190		-58		75		80		70		3		160		140		6		240		210		9
		200East		60		58		-60		40		80		3		160		160		3		160		160		9
		300West		190		-58		75		50		50		3		100		160		6		150		240		9
		300East		40		58		-60		80		50		3		160		100		6		240		150		9
Ag		100West		190		-58		75		120		110		6		240		220		12		240		220		12
		100East		60		58		-60		110		80		3		220		160		6		220		160		6
		200West		190		-58		75		80		70		3		160		140		6		160		140		6
		200East		60		58		-60		85		50		3		170		100		6		170		100		6
		300West		190		-58		75		90		45		8		180		90		16		180		90		16
		300East		60		58		-60		70		45		3		140		90		6		140		90		6

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Table 14-22: 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 14-23: 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 14-24: Search Neighbourhoods 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

14.8

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 14-7 and Figure 14-8.

LOGO

Figure 14-7: Cross-Section 425,775E Comparing Blocks Populated

with Gold Grades and Informing Data

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LOGO

Figure 14-8: Cross-Section 425, 335E Comparing Blocks Populated With Gold Grades and

Informing Data in the Intrepid Zone

SRK has also undertaken a comparative analysis between ordinary Kriging estimates and that estimated using an inverse distance (power of 2) estimator. Both estimation techniques show comparative global results.

14.9

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 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.

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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 14-25.

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 14-25: 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 Composites per Borehole		2

14.10

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.”

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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 14-26. 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 14-27.

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 an attempt to evaluate mineral reserves. Mineral reserves can only be estimated based on the results of an economic evaluation as part of a 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 14-26: Conceptual Assumptions Considered for Open Pit Resource Reporting


Parameter		Assumption
Pit Wall Angle		Average 48º
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 $1,400/oz. gold and USD $24.00/oz. silver
Mining Dilution and Losses		10.0%

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Table 14-27: 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 $1,400/oz. gold and USD $24.00/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 Project, SRK considered a conceptual pit shell defined using a gold price of USD $1,400 and a silver price of USD $24 per ounce. Following review of optimization results, SRK subdivided the block model into two (2) areas for reporting mineral resources (shown in Figure 14-9):

1.	Open pit, Measured, Indicated, and Inferred mineral resources reported inside the USD $1,400 conceptual pit shell above an elevation of -150 masl;

2.	Underground, Measured, Indicated and Inferred mineral resources are reported below an elevation of -150 masl and outside the conceptual pit shell above the elevation of -150 masl.

The resources at the Intrepid Zone are reported as underground material with an assigned Indicated and Inferred classification. Generally, an Indicated classification was assigned to those blocks that were coded in the first pass, which used a minimum of three (3) boreholes. Final manual smoothing was applied to the automatic classification to ensure the continuity of similar class blocks. The remaining blocks were classified as Inferred as they lack sufficient continuity to meet the Indicated criteria.

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LOGO

Figure 14-9: Schematic Vertical Section Illustrating Criteria Considered for Preparing the Mineral

Resource Statement for the Rainy River 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 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 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 0.30 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 gold price of USD $1,400 per ounce and a gold, metallurgical recovery of 88% and 90% for open pit and underground mineral resources, respectively, without considering revenues from other metals. The consolidated Mineral Resource Statement for the Rainy River Project gold zones is summarized in Table 14-28 and excludes gold mineralization from Zone 34 and the Silver Zone. The effective date of this ninth (9^th) Mineral Resource Statement prepared for the Rainy River Project is November 2, 2013.

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Table 14-28: Consolidated Mineral Resource Statement*, Rainy River Project, Ontario,

SRK Consulting (Canada) Inc., November 2, 2013^1,2,3,4,5


	Quantity ‘000t		Grade				Metal
Category	Quantity ‘000t		Au gpt		Ag gpt		Au ‘000 oz		Ag ‘000 oz
Direct Processing Material
Open Pit²
Measured		20,282		1.45		1.93		947		1,261
Indicated		80,411		1.35		2.55		3,486		6,584
O/P Measured & Indicated		100,693		1.37		2.42		4,433		7,846
Inferred		9,388		0.97		2.28		292		687
Underground²
Measured		89		4.95		2.75		14		8
Indicated		5,469		4.53		11.34		796		1,994
Measured & Indicated		5,558		4.53		11.20		810		2,002
Inferred		2,641		4.46		8.30		379		707
Stockpile Material³
Open Pit
Measured		6,294		0.37		1.29		74		262
Indicated		64,816		0.44		2.17		919		4,526
Measured & Indicated		71,110		0.43		2.09		993		4,788
Inferred		8,626		0.37		1.16		102		323
Combined Direct Processing and Stockpile Mineral Resources
Measured		26,665		1.21		1.79		1,035		1,531
Indicated		150,696		1.07		2.70		5,202		13,104
Measured and Indicated		177,361		1.09		2.57		6,236		14,635
Inferred		20,655		1.16		2.58		773		1,717

¹	Mineral resources are reported in relation to conceptual pit shells which are limited to 150m below sea level and are inclusive of the Intrepid Zone.

³	Direct processing material is defined as mineralization above a cut-off of 0.45 g/t gold and likely to be mined and processed directly.

⁴

Stockpile material includes all material within conceptual open pit shells above a cut-off of 0.30 g/t gold and below a 0.45 g/t gold cut-off as well as material within the CAP Zone (code 500) that is suitable for stockpiling and future processing based on average metallurgical recoveries of 88% gold and 75% silver.

⁵	Qualified Persons – The mineral resource statement was prepared by Dorota El-Rassi, P.Eng. (APEO #100012348) and Glen Cole (APGO #1416) from SRK Consulting (Canada) Inc., both Independent “Qualified Persons” as that term is defined in Canadian National Instrument 43-101.

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Mineral resources for Zone 34 and the Silver Zone are reported in Table 14-29 and Table 14-30, respectively.

Table 14-29: Mineral Resources¹ for Zone 34 (Domain 200), Rainy River Project, Ontario, SRK

Consulting (Canada) Inc., November 2, 2013


			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 Resources²
Indicated		191		0.60		0.23		0.62		1,656		1,424		3.71		1.39		3.80		317		272

^1.

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.

^2.	Open pit mineral resources are reported at a cut-off grade of 0.30 g/t gold. Cut-off grades are based on a price of USD $1,400 per ounce of gold, exchange rate of CAD $1.10 to USD $1.00, and gold recovery of 88%.

Table 14-30: Mineral Resources¹ for the Silver Zone (Domain 901), Rainy River Project,

SRK Consulting (Canada) Inc., November 2, 2013

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 Resources²

Indicated

2,108

0.54

20.34

0.89

36.51

1,378

60.49

^1.

^2.

Open pit mineral resources are reported at a cut-off grade of 0.30 g/t gold equivalent. Gold equivalent grade is based on a gold price of USD $1,400 per ounce, a silver price of USD $24.00 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 14-31 and Table 14-32 for open pit and underground areas, respectively.

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Table 14-31: Open Pit Mineral Resources1, Rainy River Project, Ontario, SRK

Consulting (Canada) Inc., November 2, 2013


	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		22,258		1.20		860		94,639		1.06		3,228		50		0.56		1
433		4,318		1.16		161		13,256		1.06		451		257		0.66		5
HS								13,774		0.66		292		4,928		0.63		100
CAP								23,558		0.57		435		217		0.54		4
Western														1,400		1.26		57
601														1,753		0.74		42
602														2,302		0.72		53
603														1,884		0.71		43
604														4,382		0.52		74
605														841		0.57		15

Total		26,576		1.19		1,021		145,227		0.94		4,406		18,014		0.68		394

^1.	Reported at a cut-off grade of 0.30 g/t gold based on a gold price of USD $1,400 per ounce of gold and assuming a gold metallurgical recovery of 90%. Other metals not considered.

Table 14-32: Underground Mineral Resources¹, Rainy River Project, Ontario,

SRK Consulting (Canada) Inc., November 2, 2013


	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		3,915		4.51		568		171		3.96		22
433								369		4.43		53		65		4.39		9
HS								33		5.50		6		129		3.59		15
CAP								13		3.46		1		412		3.33		44
Western														206		5.79		38
Intrepid								1,139		4.58		168		136		5.22		23
601														570		6.15		113
602														110		3.64		13
603														438		4.04		57
604														333		3.58		38
605														69		3.08		7

Total		88		4.97		14		5,469		4.53		796		2,641		4.46		379

^1.	Reported at a cut-off grade of 2.50 g/t gold based on a gold price of USD $1,400 per ounce of gold and assuming a gold metallurgical recovery of 90%. Other metals not considered.

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14.11

Grade Sensitivity Analysis

The mineral resources of the Rainy River Project are highly sensitive to the selection of a reporting cut-off grade. To illustrate this sensitivity, grade tonnage curves for the entire mineral resource model are presented in Figure 14-10, block model quantities and grade estimates are also presented in Table 14-33 at various cut-off grades.

Table 14-34 and Table 14-35 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.

LOGO

Figure 14-10: Rainy River Project Global Grade Tonnage Curves

(Open Pit and Underground Material Combined)

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Table 14-33: Global Block Model Quantities and Grade Estimates¹ 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		43,066		0.83		1144		547,195		0.71		12,451		1,206,710		0.24		9,367
0.2		3,884		1.01		1101		384,719		0.94		11,617		393,279		0.46		5,813
0.3		7,126		1.20		1047		263,277		1.23		10,441		214,434		0.64		4,406
0.35		24,465		1.30		1019		220,328		1.40		9,886		172,427		0.71		3,956
0.4		22,440		1.38		995		188,412		1.55		9,413		129,376		0.82		3,421
0.5		19,153		1.54		948		153,399		1.78		8,802		70,419		1.13		2,561
0.6		16,470		1.70		900		129,794		1.99		8,309		47,711		1.40		2,148
0.7		14,197		1.87		853		104,359		2.30		7,714		34,083		1.69		1,848
0.8		12,235		2.05		806		85,843		2.60		7,186		23,950		2.06		1,586
0.9		10,531		2.24		759		72,175		2.90		6,719		19,280		2.33		1,443
1.0		9,255		2.42		720		61,480		3.19		6,307		16,075		2.58		1,331
1.2		7,240		2.79		649		51,175		3.52		5,788		12,150		2.99		1,168
1.4		5,700		3.19		585		39,949		4.07		5,224		9,843		3.34		1,056
1.6		4,736		3.54		539		32,061		4.61		4,751		7,823		3.75		943
1.8		3,904		3.93		494		26,195		5.16		4,345		5,776		4.40		817
2.0		3,305		4.30		457		21,617		5.72		3,974		4,811		4.83		748
2.2		2,785		4.72		422		18,306		6.22		3,659		4,099		5.22		688
2.4		2,421		5.08		395		15,655		6.71		3,377		3,396		5.73		626
2.5		2,281		5.24		384		13,951		7.13		3,199		3,113		5.98		598
2.6		2,172		5.38		375		12,986		7.37		3,077		2,876		6.23		577
2.8		1,937		5.70		355		11,788		7.64		2,897		2,493		6.69		537
3.0		1,738		6.02		336		10,367		8.07		2,691		2,216		7.08		504
3.5		1,320		6.91		293		8,539		8.47		2,325		1,598		8.36		430
4.0		1,063		7.68		262		6,666		9.24		1,980		1,216		9.59		375
5.0		766		8.90		219		4,956		10.07		1,604		731		11.97		282
10.0		195		14.41		90		2,880		10.86		1,006		113		17.31		63

1.	The reader is cautioned that the figures in this table 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 cut-off grade.

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Table 14-34: Block Model Quantities and Grade Estimates¹ 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 0.25 g/t Au
Measured		29,712		1.10		1,048		1.72		1,642
Indicated		173,409		0.83		4,653		2.27		12,657
Measured & Indicated		203,120		0.87		5,702		2.19		14,299
Inferred		23,727		0.58		444		1.57		1,198
Cut-off Grade 0.35 g/t Au
Measured		23,975		1.29		994		1.84		1,415
Indicated		122,819		1.06		4,172		2.49		9,830
Measured & Indicated		146,794		1.09		5,166		2.38		11,245
Inferred		14,109		0.78		354		1.94		879

^1.	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.

Table 14-35: Block Model Quantities and Grade Estimates¹ 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		8,550		3.74		1,029		9.62		2,646
Measured & Indicated		8,663		3.75		1,044		9.53		2,655
Inferred		3,993		3.71		477		7.38		947
Cut-off grade 3.0 g/t Au
Measured		70		5.57		12		2.96		7
Indicated		4,331		5.06		705		13.19		1,838
Measured & Indicated		4,401		5.07		717		13.03		1,844
Inferred		5,152		5.21		863		7.47		1,237

^1.	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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Figure 14-11 presents a plan showing the distribution of the open pit mineral resources relative to the USD $1,400 conceptual pit outline.

LOGO

Figure 14-11: Distribution of Open Pit Mineral Resources Relative to the Conceptual Pit Outline

14.12

Previous Mineral Resource Estimates

A comparison between the October 10, 2012 and the November 2, 2013 Mineral Resource Statements is shown in Table 14-36. The reduction in Measured and Indicated Open Pit Mineral Resources grade is primarily due to the increase in block size in 2013 (from 5 m x 5 m x 5 m to 10 m x 10 m x 10 m) and the reduction in reporting cut-off grade from 0.35 g/t gold to 0.30 g/t gold. The reduction in Inferred resources is due to a difference in mineral resource reporting methodology in relation to a conceptual pit shell. The increase in the Underground Mineral Resources is primarily due to the addition of the Intrepid Zone to the Mineral Resource Statement.

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Table 14-36: Comparison of the October 2012 and November 2013 Mineral Resource Statements


	Quantity			Grade (g/t)						Contained Metal (oz.)
Classification	(tonnes)			Gold			Silver			Gold			Silver
Open Pit
Measured		-4	%		-10	%		-6	%		-13	%		-9	%
Indicated		15	%		-12	%		-11	%		0	%		3	%
Measured & Indicated		11	%		-13	%		-9	%		-2	%		1	%
Inferred		-81	%		-6	%		-22	%		-82	%		-85	%
Underground
Measured		1	%		0	%		0	%		0	%		0	%
Indicated		32	%		1	%		85	%		33	%		144	%
Measured & Indicated		31	%		1	%		85	%		32	%		143	%
Inferred		194	%		7	%		79	%		216	%		428	%

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15.

MINERAL RESERVE ESTIMATE

15.1

Introduction

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.

15.2

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 at the Rainy River Project will follow the standard practice of an open pit operation, with a conventional drill and blast, load and haul cycles 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.

15.2.1

Resource Block Model

The mining engineering work required for the Feasibility 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 August 6th, 2013. The model was transferred by BBA into the MineSight mining software for the open pit portion of the Feasibility Study. The unit block size in the model is 10 m x 10 m x 10 m. BBA used the same digitized topographical and bedrock mapping data provided by Rainy River for the 2013 Feasibility Study. The UTM NAD83 coordinate system was used.

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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: 1 = Measured, 2 = Indicated or 3 = Inferred);

•

Pt (platinum, not loaded in MineSight);

•

Pd (palladium, not loaded in MineSight);

•

Ni (nickel, not loaded in MineSight);

•

Cu (copper, not loaded in MineSight);

•

S (sulphur, not loaded in MineSight);

•

Ca (calcium, not loaded in MineSight);

•

AP (acid generation classification – acid potential);

•

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 and confirmed in order to insure conformity with the results provided by SRK.

15.2.1.1

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 NAD83 coordinate system as the block model:

•

Topography surface; and

•

Overburden surface (interface bedrock/overburden).

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The two (2) files are used to provide information about the percentage of overburden and rock below the topographic surface and the overburden/bedrock contact.

The overburden surface that was provided is important for understanding the large variability of overburden thickness in the pit area. The overburden thickness can reach up to 50 m in the centre region of the pit as indicated in Figure 15-1.

LOGO

Figure 15-1: Isopach Mapping of Overburden Thickness

15.2.1.2

Density

The density for the mineralized blocks, as coded in the SRK Block Model, range from 2.77 t/m³to above 3.00 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.

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15.2.1.3

Model Gold and Silver Recoveries

The gold and silver mill recoveries are determined from metallurgical testwork results (presented in Section 13.15 of Chapter 13 of this report) and are coded in the Block Model and used for pit optimization. The gold and silver recovery equations derived from the recovery curves for CAP zone and Non-CAP zones.

It is important to note that the gold and silver 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, that 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.

15.2.2

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 (LG) 3D open pit optimization 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.

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15.2.2.1

Pit Optimization Parameters

The main pit optimization parameters used in the LG 3D routine are listed in Table 15-1.

The costs for the pit optimization process were based on best information available at the time of the study, including costs from the initial 2013 Feasibility Study, gold and silver mill recoveries developed during the initial phase of the Feasibility Study, costs from similar mining operations and BBA experience.

Table 15-1: Pit Optimization Parameters


Type of Activity	Unit	Values
Mining Cost of Rock¹	$/t mined	1.95
Mining Cost of Overburden¹	$/t mined	1.50
Processing Cost¹	$/t milled	8.65
General and Administration Cost¹	$/t milled	1.21
Gold Recovery (average)	%	89.9
Gold Recovery for CAP zone (average)	%	74.3
Silver Recovery (average)	%	67.1
Silver Recovery for CAP zone (average)	%	69.5
Gold Selling Price	USD/oz.	Varied by $100 increments
Silver Selling Price	USD/oz.	25
Exchange Rate	CAD/USD	1.05
Overall Pit Slope Angle	degree	Varies from 37 to 53 depending on zone
Overall Overburden Slope Angle	degree	16
Pit Limitation		RRP Property
Depth/elevation constraint	masl	-50

^1.	Pit optimization parameters differ from the final operating cost figures. The pit optimization parameters were taken as an initial iteration to determine the optimized pit shell.

The LG 3D pit optimization was run using complex slopes approach due to various pit slope angles by sectors. It is important to note that all the slopes have been reduced by 3 degrees on average from the final design specification provided by AMEC and presented in Chapter 16 in order to account for operational design factors such as ramps, geotechnical berms and benching arrangements that will be incorporated subsequently in the final engineering design process.

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In addition to the aforementioned processing and pit slope parameters, there were also various limits and constraints that were imposed as agreed upon by BBA and New Gold. These are as follows:

•

A constraint was provided from Bayfield’s Burns property to the east of the pit (New Gold owns surface rights to the Burns property while mineral rights are owned by Bayfield Ventures Corp.); and

•

A -50 masl elevation constraint was applied to the pit optimization. In accordance with the PEA Update results, New Gold decided that the option for an open pit be limited to a depth of -50 m with the underground operation 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 15-2, along with the selected theoretical pit shell.

15.2.2.2

Pit Optimization Results

Using the technical and economic parameters described previously, the MineSight LG 3D pit optimizer tool was used to produce an optimum pit shell at various gold prices for the Rainy River deposit. Once the series of pit shells are generated, the total material moved, total in-pit resource and stripping ratios were used to select the optimum pit shell. The selection methodology was based on the stripping ratio and a mine life of approximately 15 years to maximize the NPV. Based on this analysis, the chosen optimized pit for this Feasibility Study was the pit having a gold price of US$800/oz.

A plan view of the LG 3D pit shell is shown in Figure 15-3. The pit edge commences at an elevation of approximately 350 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 15.2.2.1.

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LOGO

Figure 15-2: Rainy River Theoretical Pit Shell and Bayfield constraint (Plan View)

15.2.2.3

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 the grade required to cover the costs for processing, General and Administrative (“G&A”) costs 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 5%.

Table 15-2 presents the results of a sensitivity analysis of the gold mill COG versus various gold prices and dilution. This analysis confirms the selected COG of 0.30 g/t Au, and benchmarks well with similar gold operations.

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Table 15-2: Mill COG Calculated at Various % Dilution and Gold Prices¹


Gold Price	% Dilution using 0.2g/t Au grade
(USD/oz.)	0%		5%		10%		15%
800		0.42		0.44		0.46		0.48
900		0.37		0.39		0.41		0.42
1000		0.33		0.35		0.37		0.38
1100		0.30		0.32		0.33		0.35
1200		0.28		0.29		0.31		0.32
1300		0.26		0.27		0.28		0.29
1400		0.24		0.25		0.26		0.27

15.2.2.4

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:


		Eq Au (g/t) = Au (g/t) +		Ag (g/t) * Ag price ($/oz.) * Ag mill recovery (%)
				Au price ($/oz.) * Au mill recovery (%)

15.2.3

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 15-3.

Table 15-3: Detailed Open Pit Mine Design Parameters


Parameter		Value
Benching Arrangement		3 x 10 m
Berm Width		10.5 m – 12 m (AMEC 2013F)
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%

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The in-pit haulage roads are 33 m wide to accommodate the proposed 220-tonne haul trucks. This ramp will provide sufficient 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.

There are two (2) in-pit haulage ramps in the final pit design as follows:

•

The main ramp exits on the east side of the pit in order to provide easy access to the site infrastructure such as the primary crusher, the east rock pile and to the low grade ore stockpile;

•

The second ramp exits on the west side of the pit for a shortened access to the west rock pile and overburden pile. The designed pit is approximately 1,700 m in length by 1,400 m wide and 400 m deep. The lowest bench is at an elevation of 50 m below sea level.

Figure 15-3 presents a detailed plan view of the proposed open pit mine (final pit) and Figure 15-4 presents an isometric view of the final pit and the selected optimized pit shell. Figures 15-5 to 15-12 present typical bench plans and cross-sections of the detailed pit versus the optimized pit. A COG of 0.30 g/t Au eq is shown in the coloured blocks.

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LOGO

Figure 15-3: Detailed Open Pit Mine Design (Plan View)

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LOGO

Figure 15-4: Final Pit Design and LG Optimization – Isometric View

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LOGO

Figure 15-5: Final Pit Design and LG Optimization – Elevation 290 masl

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LOGO

Figure 15-6: Final Pit Design and LG Optimization – Elevation 160 masl

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LOGO

Figure 15-7: Final Pit Design and LG Optimization – Elevation 10 masl

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LOGO

Figure 15-8: Final Pit Design and LG Optimization – Cross Section (East 425 400, looking West)

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LOGO

Figure 15-9: Final Pit Design and LG Optimization – Cross Section (East 425 500, looking West)

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LOGO

Figure 15-10: Final Pit Design and LG Optimization – Cross Section (East 425 600, looking West)

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LOGO

Figure 15-11: Final Pit Design and LG Optimization – Cross Section (East 425 700, looking West)

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LOGO

Figure 15-12: Final Pit Design and LG Optimization – Cross Section (East 425 800, looking West)

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15.2.4

In-Pit Dilution and Mine Recovery

The mining dilution encountered during mining operation is defined as the addition of the material below the COG delivered to the mill due to over-mining at the ore/waste contacts and ore blocks below the COG inside mineralized zones that cannot be physically separated from mining methods. “Orphan” ore blocks that cannot be separated from waste are not mined. The mine recovery is the percentage of recovered ore blocks above the COG.

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.

The method used to estimate the mining dilution and mine recovery is based on a manual procedure that involves drawing mining polygons on a series of equally-spaced bench plans inside the mine design area. This method is based on the following guidelines to ensure that the estimate is consistent and systematic throughout the selected benches:

•

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);

•

The “orphan” blocks are not mined; and

•

The estimation of the in-pit dilution was carried out on five (5) selected equally spaced benches.

Using the assumptions outlined above, BBA has estimated a total mining dilution of 7.79% at a grade of 0.21 g/t Au in the 10m x 10m x 10m block model for the Feasibility Study.

In the 5m x 5m x 5m block model for the original 2013 Feasibility Study, the average grade was 0.998 g/t Au, based on a COG of 0.30 g/t Au. Using the same COG, the average grade became 0.955 g/t Au in the 10m x 10m x10m block model for the Feasibility Study, indicating that a “built-in” dilution was inherently added when using larger unit block sizes. BBA has estimated the “built-in” dilution in the block model and the net mining dilution as follows:

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• Total dilution estimate in the 10m x10m x 10m model:		7.79%
• Gold average grade in the 5m x 5m x 5m model:		0.998 g/t
• Gold average grade in the 10m x 10m x 10m model:		0.955 g/t
• “Built-in” dilution in 10m x 10m x 10m block model:		4.31%,e.g.(0.998-0.955)/0.988
• Net dilution estimate in the 10m x 10m x 10m model:		3.48%, e.g. 7.79% - 4.31%

Table 15-4 presents the final values of dilution, dilution grade and mine recovery (or ore loss) used.

Table 15-4: Estimation of In-pit Dilution and Mine Recovery


Parameter	Units	Value
In-Pit Dilution	%	4.0
Dilution Au Grade	g/t Au	0.21
Dilution Ag Grade	g/t Ag	1.19
Mine Recovery (ore loss)	%	95 (5)
15.2.5 Open Pit Mineral Reserves

15.2.5

Open Pit Mineral Reverves

Using the engineered pit design, the open pit mine contains 100.1 Mt of diluted reserves in the Proven and Probable categories at an average grade of 0.96 g/t Au and 2.49 g/t Ag. Total waste material amounts to 318.2 Mt of waste rock and 73.6 Mt of overburden resulting in an overall open pit strip ratio of 3.91 (tonnes of waste rock and overburden per tonne of ore). Table 15-5 presents the final open pit Mineral Reserves for the Rainy River pit.

Table 15-5: Open Pit Mineral Reserves Statement^1,2


Open Pit Mineral Reserves	Tonnage (Mt)		Gold Grade (g/t)		Silver Grade (g/t)		Contained Gold (oz)		Contained Silver (oz)
Proven		22.7		1.14		1.88		830 279		137 0407
Probable		77.4		0.91		2.67		2 274 646		6 651 985

TOTAL		100.1		0.96		2.49		3 104 924		8,022,391

^1.

Open pit mineral reserves have been estimated using a cut-off grade of 0.30 g/t gold-equivalent, based on metal prices of USD $800 per ounce gold and USD $25 per ounce silver, a foreign exchange rate of CAD$1.05 to USD$1.00, gold recovery of 89.9% (non-CAP zone) and 74.3% (CAP zone) and a silver recovery of 67.1% (non-CAP zone) and 69.5% (CAP zone).

^2.	Open pit reserves have been estimated using a dilution of 4% at 0.21 g/t Au and 1.19 g/t Ag. Open Pit reserves have been estimated using a mine recovery of 95%.

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15.3

Underground Mining

The underground mine design and the underground Mineral Reserve estimate were completed by AMC to a level of detail normally associated with feasibility studies. The Mineral Reserve estimate stated herein is consistent with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Mineral Reserves and is suitable for public reporting.

Inferred resources within in the underground mine plan account for 0.07% of the Mineral Reserves tonnage and were assigned zero grade values. AMC considers this inclusion immaterial.

The Mineral Reserves were developed from two (2) independent resource models that cover resources beneath the open pit (“ALL_FP1.dm” block model) and those in the Intrepid zone (“ALL_FINAL2.dm” block model). The two (2) resource models were provided by SRK to AMC on behalf of New Gold in August 2013 and September 2013 respectively.

15.3.1

Underground Mineral Reserves

15.3.1.1 Orebody Description

Ore-grade mineralization occurs in seven independent areas of differing geometry and grade; ODM West, ODM Main (Upper and Lower zones), ODM East, 433, 17 East Lower, 17 East Upper and the Intrepid zone. The bulk of the underground Mineral Reserves are contained within the ODM Main and Intrepid Zones (refer to Figure 15-13).

Ore occurs in subvertical horizons in varying widths from about 3 m to 20 m. Widths over 15 m are rare, and the weighted average across all zones is approximately 8 m. The ore footwall (FW) and hangingwall (HW) generally dip at 60 degrees or more, but can flatten locally to as low as 45 degrees in some areas. Figure 15-14 and Figure 15-15 illustrate typical sections through the ODM Main and Intrepid zones respectively.

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LOGO

Figure 15-13 : Isometric View of the Rainy River Underground Mine

LOGO

Figure 15-14 : Typical ODM Main Zone Cross-Sections

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LOGO

Figure 15-15 : Typical Intrepid Zone Cross-Sections

15.3.1.2 Cut-off Grade (“COG”)

A preliminary estimate of breakeven operating costs was made early in the Feasibility Study to define mining shapes from which the Mineral Reserves would be formed.

It was recognized at the time that it might be appropriate to lower the COG applied in the initial 2013 Feasibility Study (3.5 g/t Au eq) to 2.5 g/t Au eq based on the preliminary estimate of site operating costs (Table 15-6).

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Table 15-6: Preliminary Estimation of Operating Costs and Breakeven Cut-off Grade


Parameter	Units	Values
Mining Cost	$/t		75.52
Processing Cost	$/t		8.65
General and Administration Cost	$/t		1.21
Refining & Transport	$/t		0.14
Sustaining Capital	$/t		2.00
Royalties	$/t		0.54
Total	$/t		88.06
Gold Recovery	%		88	%
Silver Recovery	%		75	%
Gold Selling Price	$USD/oz		1,250
Silver Selling Price	$USD/oz		20
Exchange Rate	CAD:USD		1.0
Breakeven Au equivalent (Au eq)	g/t		2.49

The costs shown above were largely drawn from the initial 2013 Feasibility Study, resulting in a potential breakeven COG of 2.49 g/t Au eq. However, AMC recognized the possibility of an increase in the site operating cost projection during the Feasibility Study. An increase in mining costs to $100/t was considered to be a reasonable expectation given that the previous mining plan incorporated a significant amount of Cut-and-Fill (“C&F”) mining. The potential impact would be to increase site operating costs to $112/t and result in a breakeven COG of 3.2 g/t Au eq. Any potential increase in surface costs would further increase the breakeven COG.

Considering the above, the 3.5 g/t Au eq COG was retained for the purpose of the Feasibility Study. This COG underpins the current statement of underground Mineral Reserves.

All inputs into the calculation of the breakeven COG evolved over the course of the Feasibility Study, including the metal price scenario, metallurgical recoveries, exchange rate, underground mining costs and surface costs. Estimation of the breakeven COG based on the final inputs to the Feasibility Study yields a value of 2.75 g/t Au eq (Table 15-7).

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Table 15-7: Updated Estimation of Breakeven Cost and COG


Parameter	Units	Values
Mining Cost	$/t		90.10
Processing Cost	$/t		9.00
General and Administration Cost	$/t		1.50
Refining & Transport	$/t		0.14
Sustaining Capital	$/t		6.71
Royalties	$/t		2.37
Total	$/t		109.82
Gold Recovery	%		95	%
Silver Recovery	%		75	%
Gold Selling Price	$USD/oz		1,300
Silver Selling Price	$USD/oz		22
Exchange Rate	CAD:USD		1.05
Breakeven Au equivalent (Au eq)1	g/t		2.75

1.	Also includes recognition of backfill dilution

The results indicate that COG optimization work should be undertaken at a later stage. Lowering the COG may not necessarily increase project value, and may negatively impact the head-grade in the early years of mine life. However, a detailed evaluation is required to understand all impacts associated with a change in COG.

The underground Mineral Reserves include material from both longhole stopes and the development within mineralized horizons required to access those stopes. Development through low-grade areas is inevitable in the current mine plan. As such, a lower (1.5 g/t Au eq) cut-off grade was applied to development within mineralized horizons on the basis that the mining cost is effectively sunk, and the remaining costs to process this material as mill-feed are marginal. Material below the 1.5 g/t Au eq COG will be retained underground as backfill.

15.3.1.3 Gold Equivalent (Au eq) Calculation

The application of a metal equivalent COG considered both payable metals, gold and silver. The gold to silver ratio varies throughout the deposit, most notably in the Intrepid Zone, that contains significantly more silver than the other zones. Metal prices and process recoveries were provided by New Gold.

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The Mineral Reserves have been defined on a 3.5 g/t Au eq COG. The Au eq value of each block in the resource block models was assigned through the expression below, using the parameters shown in Table 15-8:


		Eq Au (g/t) = Au (g/t) +		Ag (g/t) * Ag price ($/oz.) * Ag mill recovery (%)
				Au price ($/oz.) * Au mill recovery (%)

Table 15-8: Au eq Calculation Parameters


Parameter	Units	Value
Gold Selling Price	$USD/oz	1,250
Gold Recovery	%	88
Silver Selling Price	$USD/oz	20
Silver Recovery	%	75

15.3.2

Mining Shapes

AMC used the Mineable Shape Optimizer (“MSO”) module from the Datamine Studio 3 mine planning software package to produce design excavations (shapes) that meet both the COG and operational design criteria.

The design criteria constrain the geometry of all planned excavations to what is achievable through the planned mining methods. The preliminary shapes were refined to minimize the amount of sub-economic material within the shape volumes. Chapter 16 provides further detail on mining shapes and design parameters.

In numerous cases, the MSO process identified mineable shapes around small tonnages of mineralization meeting COG but at a considerable distance from the nearest centre of mining activity. These shapes (orphans) were invariably removed as being uneconomic to mine after consideration of the development cost required to access them.

15.3.3

Dilution and Recovery Estimates

The Mineral Reserves account for planned and unplanned dilution.

Planned dilution represents the portion of the design excavation (shape) that is sub-economic on a standalone basis, but must be mined to render the economic portion recoverable for practical reasons.

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Unplanned dilution includes those sources of sub-economic material that are external to the design excavation but inevitably dilute the ore in varying degrees. Unplanned dilution is manageable through proper design and mining practice, but is inherent to all operations and must be added to the base tonnage within the design mining shapes in the estimation of Mineral Reserves. These sources include waste rock dilution from stope wall sloughing and/or blasting over-break, backfill dilution from sloughing of the cemented aggregate fill (“CAF”) exposure of the previous stope in the sequence, and over-mucking of backfill that will often form the stope floor.

Stope sloughing may occur at the stope walls, and particularly at the hangingwall. The equivalent linear over-break/slough (“ELOS”) method was used by AMEC in the estimation of slough volume for the 2013 Feasibility Study. The results indicated that, on average, 0.25 m of HW overbreak/slough can be expected. Likewise, 0.25 m footwall dilution on average was estimated due to blasting over-break. AMC considers the AMEC overbreak/slough dilution assessment to be reasonable. The results are illustrated in Figure 15-16.

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Figure 15-16 : Unplanned Dilution from Hangingwall and Footwall Rock

The total unplanned dilution from rock sources (hangingwall and footwall) is estimated to be 7.4% of the design stope tonnage, as shown in Table 15-9. The design stope shapes were expanded by 0.25 m on both the hangingwall and footwall sides such that the metal value in the incremental volume would be captured when the shapes were evaluated against the block models.

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Table 15-9: Dilution Estimate for Hangingwall Slough and Blasting Over-break


Dilution source	Depth (m)		Percent (%)
Hanging wall slough dilution		0.25		3.7
Blasting overbreak dilution		0.25		3.7
		Total		7.4

Backfill dilution will originate from two sources: over-mucking the floors in the longitudinal longhole open stoping (“LHOS”), and sloughing of cemented rock fill (“CRF”) from the end of the longitudinal LHOS. Backfill dilution is assumed to carry no metal grades in the compilation of Mineral Reserves.

In the longitudinal LHOS, the mucking will be a combination of manual and remote mucking. Manual mucking provides a high degree of operator control and can minimize the introduction of dilution into the production stream. Conversely, remote mucking results in less operator control and it is estimated that, on average, over-digging into the floor will introduce 0.25 m of backfill dilution.

Sloughing of the cemented aggregate fill from the previous stope in the sequence has been estimated at 0.75 m width on average. Figure 15-7 illustrates the two sources of backfill dilution.

LOGO

Figure 15-17 : Backfill Dilution

The total estimated stope dilution from end-wall sloughing and over-mucking the floor has been summarized in Table 15-10. End-wall sloughing introduces 3.3 % dilution and over-mucking of the stope floor introduces 1.0 % dilution.

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Table 15-10: Dilution Estimate from Backfill Sources


Dilution Source	Depth (m)		Percent (%)
End-wall sloughing from CAF		0.75		3.3
Over-mucking from rockfill and CAF		0.25		1.0
		Total		4.3

15.3.4

Recovery Factors

Mining recovery factors are applied to account for the percentage of ore in the design stope that is not recovered. With longitudinal LHOS, potential ore losses can occur due to the difficulty in mucking in stope corners and edges, inefficient drilling and/or blasting in stope corners and walls, or abandoning of stopes due to excessive dilution from waste rock or backfill. AMC has applied 95% recovery to the stope inventory based on its experience.

15.3.5

Mineral Reserves

The total unplanned dilution is estimated to be 11.7% of the stoping inventory based on 7.4% rock dilution and 4.3% backfill dilution. Recovery of the stoping inventory is estimated to be 95%. Ore development is considered to be 100% recoverable with no dilution. When dilution is expressed on a global basis, inclusive of ore development tonnage, the average dilution of the Mineral Reserves is 8.3% and the average recovery is 96.5%. Table 15-11 presents the final underground Mineral Reserves across all zones.

Table 15-11: Underground Mineral Reserves Statement¹


Underground Mineral Reserves	Tonnage (‘000 t)	Gold Grade (g/t)	Silver Grade (g/t)	Contained Gold (‘000 oz)	Contained Silver (‘000 oz)
Probable	4,187	4.96	10.31	668	1,388

Underground Mineral Reserves include 3.29 Mt of ore above a 3.5 g/t Au eq COG, grading 5.58 g/t Au and 10.72 g/t Ag, and a further 0.89 Mt of ore grading 2.70 g/t Au and 8.81 g/t Ag between 1.5 g/t Au eq COG and 3.5 g/t Au eq COG. Au price and metallurgical recovery assumed to be $USD 1,250/troy ounce and 88% respectively. Ag price and metallurgical recovery assumed to be $USD 20/troy ounce and 75% respectively. Exchange rate of $CAD = $USD. Average dilution and mining recovery estimated at 8.3% and 96.5% respectively.

15.4

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 15-12 shows a detailed summary of the in-pit and underground reserves.

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Table 15-12: Open Pit and Underground Proven and Probable Mineral Reserves^1,2,3,4,5,6


Mineral Reserves Category	Tonnage (‘000 t)		Au Grade (g/t)		Ag Grade (g/t)		Contained Metal Au (koz)		Contained Metal Ag (koz)
Direct Processing Material
Open Pit
Proven		15,839		1.47		2.04		746		1,038
Probable		46,866		1.26		3.05		1,896		4,594
Underground
Proven		—		—		—		—		—
Probable		4,187		4.96		10.31		668		1,388

Total Direct Processing Material		66,892		1.54		3.26		3,311		7,021

Stockpile Material
Open Pit
Proven		6,843		0.38		1.51		84		332
Probable		30,541		0.39		2.10		378		2,058

Total Stockpile Material		37,384		0.38		1.99		462		2,390

Combined Direct Processing and Stockpile Material
Open Pit
Proven		22,681		1.14		1.88		830		1,370
Probable		77,407		0.91		2.67		2,275		6,652

Total		100,088		0.96		2.49		3,105		8,022

Underground
Proven		—		—		—		—		—
Probable		4,187		4.96		10.31		668		1,388

Total		4,187		4.96		10.31		668		1,388

Total Combined
Proven		22,681		1.14		1.88		830		1,370
Probable		81,594		1.12		3.06		2,943		8,040

TOTAL		104,275		1.13		2.81		3,773		9,410

^1.

Open pit mineral reserves have been estimated using an optimized pit shell based on metal prices of USD $800 per ounce gold and USD $25 per ounce silver, a foreign exchange rate of CAD $1.05 to USD $1.00, gold recovery of 89.9% (non-CAP zone) and 74.3% (CAP zone) and a silver recovery of 67.1% (non-CAP zone) and 69.5% (CAP zone). The cut-off grade is based on a gold price of USD $1,200. Underground reserves have been estimated from mining shapes generated using a cut-off grade of 3.5 g/t gold-equivalent. Development material from stope access drives above a cut-off grade of 1.5 g/t gold-equivalent is also assumed to be sent to the mill for processing. Underground breakeven cut-off grade calculated at 2.75 g/t gold-equivalent based on metal prices of USD $1,300 per ounce gold and USD $22 per ounce silver, a foreign exchange rate of CAD $1.05 to USD $1.00, gold recovery of 95% and a silver recovery of 75%.

^2.

Open pit reserves have been estimated using a dilution of 4% at 0.21 g/t Au and 1.19 g/t Ag. An average dilution of 11.7% for the underground stoping (8.3% total underground, inclusive of development ore that includes dilution from both overbreak and backfill. Open pit and underground reserves have been estimated using a mine recovery of 95% and 96.5%, respectively.

^3.	Open pit direct processing material is defined as mineralization likely to be mined and processed directly and above a variable cut-off grade ranging from 0.3-0.7 g/t.

^4.	Stockpile material includes all material within designed open pit between variable cut-offs described above in Note 3, as well as material within the CAP zone (code 500) that is suitable for stockpiling and future processing.

^5.

Mineral Reserves for the open pit are derived from the resource model effective November 2, 2013. Models for the underground reserves were derived from the August 2013 and September 2013 models for the main ODM zone and Intrepid Zone, respectively. Models were 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 Canadian National Instrument 43-101. The combined mineral resource statement, including the Intrepid Zone was provided to BBA on November 2, 2013. Rainy River’s exploration program in Richardson Township is being supervised by Mark A. Petersen, (AIPG Certified Professional Geologist #10563), Vice President, Exploration for New Gold and a “Qualified Person” as defined in Canadian National Instrument 43-101. New Gold continues to implement a rigorous QA/QC program to ensure best practices in drill core sampling, analysis and data management.

^6.

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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 reserve extraction. The milling throughput rate was established at 21,000 tpd (19,500 tpd from the open pit and 1,500 tpd from underground when full production is achieved). The location of the open pit, underground access ramp (main portal), haul roads, stockpiles and waste rock piles are shown on the Project general layout in Appendix F. The open pit operations will deliver material to a gyratory crusher for primary size reduction and delivery to the processing plant. Ore from the underground operation will be fed to a portable crushing system for size reduction and tramp metal removal prior to being delivered to the gyratory crusher. Utilization of New Gold’s mining equipment and personnel is envisaged for the development of the open pit, as well as for the removal of overburden. In addition, a combination of New Gold personnel and various contractors will be used for underground development and ongoing production activities.

Figure 16-1 shows a representation of the underground and open pit mine at the end of Year 2030, looking north.

LOGO

Figure 16-1: Isometric View of the Rainy River Open Pit and Underground Mine, Looking North

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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 1,700 m [length] x 1,400 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°. For the most part, the bedrock 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, 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 that 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 dip towards the west at 85°. A persistent east dipping fault striking north-south transects the pit (Figure 16-2); 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 joint 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 specimen 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 (the predominant wall rock) were found to be strong, to very strong rocks ranging from 71 to 116 MPa in UCS.

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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 (“BFA”) will be cut at 70° to 75°, with inter-ramp angles (“IRA”) comprised between 54.5° and 56.3° and overall slope angles (“OSA”) in bedrock of 40.6° to 54.5°, 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 an 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 (°)	Bench Height (m)	Bench Width (m)	IRA (°)
FLS	5, 10	50	3x10	10.5	40.0
FLS	3	55	3x10	10.5	43.6
FLS	4, 6, 7, 9	70	3x10	10.5	54.5
FLS	1	73	3x10	12	54.8
FLS	2, 8	75	3x10	12	56.3

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).

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LOGO

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

16.2.1.2

Waste Rock and Overburden Stockpile Design

Geotechnical slope stability recommendations were provided for the mine waste rock and overburden stockpiles (AMEC 2013c). The rate of stockpile development will control stability due to excess pore pressures generated in the foundation soils (AMEC 2013a). From a stability perspective, the slopes at the south perimeter of the West Mine Rock Stockpile and west side of the East Mine Rock Stockpile are critical, as the height is greatest and the upper varved clay strata is 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 safety factors.

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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 Mine Rock 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. The EO methodology is an integrated approach to maximizing the NPV of a mining operation by simultaneously optimizing up to ten (10) different mechanisms across the mining value chain. The main objective 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 software. Prober is a Whittle Consulting in-house software used for strategic mine planning.

The EO Study was conducted to support the PEA Update and subsequent studies for the Rainy River 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 the June 2011 block model include:

•

Base case assessment;

•

Variable cut-off grade and stockpile value contributions;

•

Mining capacity trade-off;

•

Processing capacity trade-off;

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•

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 with a pit phasing and stockpile movement strategy by period that 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 Update Study. The use of an elevated cut-off grade strategy over the life of mine has resulted in a significant amount of low grade ore stockpiled material.

16.2.2.2

Open Pit Mining Phases

Based on the EO report, BBA prepared an open pit mine production schedule based on the August 2013 block model provided by SRK. In order to optimize the operational stripping ratio in the early years of the Project, and to increase the Project’s NPV, two (2) optimized shells for an initial pit (Phase 1) and an intermediate pit (Phase 2), representing four (4) to five (5) years of mining each, were generated using the Lerchs-Grossman 3D MineSight optimization algorithm. The open pit mining phases were designed using the same design parameters as the final pit design (Phase 3) and were used as a guide to prepare the first few years of the LOM mine plan. Between each mining phase, there will be a transitional period during which stripping of the next phase will be started before the end of the previous phase in order to maintain a constant supply of ore to the mill.

The mining phases presented in Figure 16-3 show the mining phases in 3D view where the years represent the rock excavation period. Figures 16-4 and 16-5 show the mining phases in longitudinal cross-section.

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LOGO

Figure 16-3 - Phase 1, Phase 2 and Final Pit - 3D View

LOGO

Figure 16-4: Phase 1, Phase 2 and Final Pit – Cross-section (East 425 500,

Looking West)

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LOGO

Figure 16-5: Phase 1, Phase 2 and Final Pit – Cross-section (East 425 700,

Looking West)

16.2.2.3

Elevated Cut-Off Grade (“COG”) Strategy

Based on the strategic mine plan study prepared by Whittle Consulting, an elevated COG was strategically employed from years 2016 to 2024 with the objective to maximize the Project’s IRR and NPV. 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 low grade ore material stockpiled during the open pit operations amounts to 37.5 Mt at an average grade of 0.39 g/t Au and 1.99 g/t Ag.

16.2.2.4

Mill Ramp-up

The Project production schedule was developed using the mill ramp-up as follows:

•

November 2016: from 0% to 30% (average 3.15 ktpd during the month);

•

December 2016: from 30% to 60% (average 9.45 ktpd during the month);

•

January 2017: from 60% to 100% (average 16.8 ktpd during the month); and

•

February 2017: 100% (21 ktpd).

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The milling rate remains at 21 ktpd until the mineral reserves are depleted.

16.2.2.4.1

Pre-Production Period

The mining production schedule is based on a pre-stripping period of 24 months, starting in Q1 2015 and ending in Q4 2016. Mill start-up is slated to begin at the end of Q4 2016. The pre-production plan was developed on a quarterly basis using the MineSight-Interactive Planner tool.

The pre-production first quarter starts in the bedrock outcrop above the CAP zone (south part of Phase 1), to provide aggregate for the construction of the plant area and access roads. The remainder of the pre-production period focus is to prepare for mine operation by opening faces on ore in Phase 1 pit. The overburden stripping of Phase 2 starts during the last quarter of the pre-production period (Q4 2016). Approximately 22.4 Mt of material is moved in 2015 (Year -2) and 25.7 Mt in 2016 (Year -1), see Figure 16-6 and Figure 16-7.

LOGO

Figure 16-6: Open Pit Mine Planning – End of 2015 (Year -2)

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LOGO

Figure 16-7: Open Pit Mine Planning—End of 2016 (Year -1)

16.2.2.4.2

Production Period

The production period plan has been designed on a quarterly basis for 2017 (Year 1) and the first two (2) quarters of 2018 (Year 2), on a semi-annual basis for the second half of 2018 (Y2) and 2019 (Year 3), and on an annual basis for the following years.

Ore production starts in November 2016 in Phase 1 and an elevated COG of 0.6 g/t Au to 0.7 g/t Au is applied for the first two (2) years of production. Phase 1 continues until the beginning of 2020 (Y4). Ore is mined from Phase 2 in 2018 (Year 3) to 2022 (Year 6) and from Phase 3 in 2021 (Y5) to 2025 (Y9).

After the mill ramp-up period (Q4 2016 – Q1 2017), 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 and 2028, the open pit mine and underground mine will feed approximately 19,500 tpd and 1,500 tpd of ore, respectively, to the mill.

The total material moved is 53.2 Mt in 2017 (Year 1) and ramps up to a maximum annual mining rate of approximately 68 Mt in 2019 (Year 3) and 2020 (Year 4). Open Pit mine production slowly decreases until the pit is depleted mid-way through 2025 (Year 10).

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Figure 16-8 to Figure 16-12 show a series of key end-of-period maps.

LOGO

Figure 16-8: Open Pit Mine Planning – End of 2017 (Year 1)

LOGO

Figure 16-9: Open Pit Mine Planning – End of 2018 (Year 2)

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LOGO

Figure 16-10: Open Pit Mine Planning – End of 2020 (Year 4)

LOGO

Figure 16-11: Open Pit Mine Planning – End of 2022 (Year 6)

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LOGO

Figure 16-12: Final Pit Contour—End of 2025 (Year 9)

16.2.2.4.3

Low-grade Ore Stockpile Rehandling Period

During the open pit operation, 37.5 Mt of low-grade ore at an average grade of 0.39 g/t Au and 1.99 g/t Ag are stockpiled in an effort to increase the initial feed grade to the mill. From the total, 4.2 Mt of ore from the CAP zone, regardless of the gold grade, will be separated and kept on a designated area on the low-grade ore stockpile for processing at the end of the mine life due to its resistance to cyanide leaching and hence, lower gold recoveries.

The low-grade ore stockpile is located between the East Mine Rock Stockpile and the primary crusher, in the East Mine Rock Stockpile area. The East Mine Rock Stockpile is approximately 1 km (road distance) from the primary crusher. After the pit is depleted, the mill is fed at a rate of 19,500 to 21,000 tpd from the low-grade ore stockpile mid-way through 2025 (Year 9) to 2030 (Year 14), depending on the underground production in Years 2025 (Year 9) to 2028 (Year 12).

The open pit material movement trends over the life of the mine can be seen graphically in Figure 16-13. The effect of the stockpiling strategy is significant as the mill head grade (green line) is higher in the early years due to a higher mill COG when compared to the average mill head grade of 0.96 g/t Au over the mine life.

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LOGO

Figure 16-13: Open Pit Material Movement over the Life-of-Mine

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16.2.3

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 site reclamation. The waste rock and overburden piles presented in this section were designed without removing any waste material that might be used for site construction. All design parameters for the various stockpiles are based on the AMEC recommendations (AMEC 2013D, E, F, G).

The waste rock material will be placed onto two (2) waste stockpiles: the West Mine Rock Stockpile (the non-potentially acid generating pile (“NPAG”)) and the East Mine Rock Stockpile (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 50% of the waste rock in the open pit is considered to be PAG material.

The West Mine Rock Stockpile area is located west of the pit and contains the NPAG and the overburden (“OB”) material. The East Mine Rock Stockpile 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 16-14.

Plan views of the stockpiling areas with slope and height details are presented in Figure 16-15 and Figure 16-16.

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Figure 16-14: Site Plan Showing West and East Mine Rock Stockpiles

LOGO

Figure 16-15: East Mine Rock Stockpile Area

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LOGO

Figure 16-16: West Mine Rock Stockpile Area

16.2.3.1

Waste Mine Rock Stockpile Design

The waste mine rock stockpiles 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. An in-situ waste rock density of 2.8 t/m³ and a swell factor of 30% have been assumed for the design of the waste mine rock stockpiles. The West and East Mine Rock Stockpiles are located and sized to fit entirely within New Gold’s land holdings and are kept at an adequate distance from all major water basins.

The waste mine rock stockpiles were designed according to the 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 on the East Mine Rock Stockpile area, as shown in Figure 16-15. Due to the bearing capacity of the ground in the south area of the East Mine Rock Stockpile, AMEC also recommends that a shallower inter-ramp angle of 8H:1V be used for the south slope of the East Mine Rock Stockpile.

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


Waste Mine Rock and Low-Grade Ore Piles	Unit	Value
Bench Face Angle	degrees	33
Inter-ramp Slope Angle	degrees	Varied from 3H:1V to 8H:1V
Catch Bench Width	m	Varied from 20 to 66
Bench Height	m	15
Ramp Width	m	33
Ramp Grade	%	10
Swell Factor	%	30

The West (NPAG) and East (PAG) Mine Rock Stockpiles have a capacity of approximately 75 Mm³ each and are built in 5 m lifts, with 15 m bench heights. Stockpiling has been sequenced in phases to allow for shorter hauls during earlier years of operation. The design summary of the West and East Mine Rock Stockpiles can be found in Table 16-3.

Table 16-3: Waste Mine Rock Stockpile Characteristics


Parameters	Unit	Value
West Mine Rock Stockpile (NPAG)
Height	m		60
Top Elevation	masl		410
Footprint Area	M m²		2.2
East Mine Rock Stockpile (PAG)
Height	m		45
Top Elevation	masl		422
Footprint Area	M m²		3.0

16.2.3.2

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 Mine Rock Stockpile next to 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 low grade ore stockpile is also located entirely within

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Rainy River’s mining claims. The low-grade ore stockpile has a capacity of 17 Mm³ and is designed according to the same design parameters as the waste rock piles in Table 16-2. The design summary of the low-grade ore stockpile can be found in Table 16-4.

Table 16-4: Low-Grade Ore Stockpile Characteristics


Low-Grade Ore Stockpile	Unit	Value
Height	m		45
Top Elevation	masl		422
Footprint Area	M m²		0.57

16.2.3.3

Overburden Stockpile Design

Overburden material will be removed during the period from 2015 (Year -2) to 2021 (Year 5). The Overburden Stockpile located west of the pit will have a capacity of 50 Mm³. The material property assumptions used for the design of the Overburden Stockpile is an in situ density of 1.8 t/m³ and a swell factor of 20%. The stockpile is entirely located within Rainy River’s mining claims or leasing claims.

The Overburden Stockpile is designed according to the design parameters presented in Table 16-5. These parameters were selected according to an inter-ramp slope of 8H:1V, as recommended by AMEC (AMEC 2013).

Table 16-5: Overburden Stockpile Design Parameters


Overburden Stockpile	Unit	Value
Bench Face Angle	degrees	20
Inter-ramp Slope Angle		8H:1V
Berm Width	m	81
Bench Height	m	15
Ramp Width	m	33
Ramp Grade	%	10
Swell Factor	%	20

The Overburden Stockpile will be built in 5 m lifts, with 15 m bench heights. The design summary can be found in Table 16-6.

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Table 16-6: Overburden Stockpile Design Summary


Overburden Stockpile	Unit	Value
Height	m		60
Top Elevation	masl		410
Footprint Area	M m²		1.95

16.2.3.4

In-Pit Dumping

In order to reduce the cycle time, 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 from the end of Year 7 to the end of Year 8 to store a portion of the NPAG material.

Figure 16-17 shows an isometric view of the in-pit dump and the dumping sequence proposed.

LOGO

Figure 16-17: In-Pit Dumping Area and Dumping Sequence

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16.2.4

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 one (1) electric hydraulic shovel. The mining fleet requirement was calculated using the production schedule presented in Table 16-27. All equipment is assumed to be owned, operated and maintained by New Gold.

Open pit mine operations are based on 720 shifts per year and correspond 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 conditions. The open pit 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, mechanical utilization, haulage distance assumptions, cycle time assumptions, truck speed and fuel consumption profiles. The primary mining fleet consists of the following:

•

The primary loading equipment for overburden, waste rock and ore are: two (2) diesel hydraulic shovels with a rated bucket capacity of 26 m³ and one (1) electric hydraulic shovel with a rated bucket capacity of 29 m³. One (1) wheel loader (18 m³ class) will be used on an as needed basis to complete the loading equipment fleet. However, the loader will be used to its full capacity from 2018 (Year 2) to 2021 (Year 5) and an extra unit (rented) will be required as well for 2019 (Year 3) and 2020 (Year 4). 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 220-tonne class payload, which matches well with the selected hydraulic shovels. The haul truck fleet starts with six (6) trucks in the pre-production phase and will gradually increase to a maximum of 22 trucks in Year 2020 (Year 4 of production); and

•

Production drilling will be accomplished using a fleet of diesel powered Down-the-Hole (“DTH”) blast hole rigs drilling 8¹/₂” diameter holes.

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16.2.4.1

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 with 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 16-7 provides information on the scheduled delays taken into consideration.

Table 16-8 shows how net operating hours (“NOH”) are derived from scheduled and unscheduled delays. Unscheduled delays are defined as delays that are outside of human control and cannot be predicted. 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.

Table 16-7: Operating Shift Parameters


Shift Parameters	Value			Unit
Shifts per 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	%

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Table 16-8: 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

16.2.4.2

Equipment Mechanical Availability, Utilization and Operator Skill Factors

For each piece of major equipment, mechanical availability, utilization and operator skill factors were designated. The mechanical availability is a percentage that represents the hours when the equipment cannot be operated due to breakdowns or planned maintenance. 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. It takes into account the low efficiency of the equipment operator during the first months of utilization (learning curve). The mechanical availability and utilization factors used over the LOM are illustrated in Table 16-9.

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Table 16-9: Major Equipment Availability and Utilization


	Time Interval
Equipment	0 - 6,000 hrs			6,000 - 12,000 hrs			12,000 - 18,000 hrs			18,000 - 24,000 hrs			24,000 - 30,000 hrs			30,000 hrs - expected life of equipment
Haul Trucks
Haul Truck Availability		88	%		88	%		87	%		87	%		86	%	83% – 85%
Haul Truck Utilization		95	%		95	%		95	%		95	%		95	%	95%
Shovels
Diesel Shovel Availability		88	%		87	%		85	%		85	%		84	%	81% – 83%
Diesel Shovel Utilization		95	%		95	%		95	%		95	%		95	%	95%
Electric Shovel Availability		89	%		88	%		86	%		86	%		85	%	81% – 83%
Electric Shovel Utilization		95	%		95	%		95	%		95	%		95	%	95%
Drills
Drill Availability		88	%		87	%		86	%		84	%		83	%	83%
Drill Utilization		95	%		95	%		95	%		95	%		95	%	95%

The operator skill factor has been applied to haul trucks, hydraulic shovels, drills and loaders, as presented in Figure 16-10.

Table 16-10: Major Equipment Operator Skill


Year	Operator Skill Factor
2015 (Year - 2)		90	%
2016 (Year - 1)		90	%
2017 (Year 1) and +		100	%

A mechanical availability of 88%, a variable utilization and an operating skill of 100% have been used for other support equipment.

16.2.4.3

Drilling and Blasting

The drill and blast design for the Study was developed by BBA, in collaboration with explosive suppliers familiar with this type of operation.

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The ore zones 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 larger drilling pattern of 7.0 m x 7.5 m. The spacing and burden for the ore zone is tighter to produce better fragmentation and selectivity. It was assumed that 20% of the waste zone will be drilled using the ore zone pattern.

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 effectiveness of the explosive column. 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.

Blasting will be conducted using a 100% emulsion-type explosive with an average density of 1.25 kg/m³. 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 0.323 kg/tonne in ore and 0.224 kg/tonne in waste.

It is also assumed that pre-split will be required for the final walls. Pre-split is executed on a 30 m bench using a spacing of 1.7 m and a hole diameter of 5.5 inches. The holes are loaded with continuous water gel or emulsion cartridges which are more precise and practical to use for this application.

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		7.0
Burden	m		6.5		7.5
Penetration rate	m/hr		42.0		42.0
Re-drill	%		10		10
Rock mass/hole	t		1,019		1,470

Bulk Emulsion
Density	kg/m³		1,250		1,250
kg/hole	kg/hole		329		329
Powder factor	kg/tonne		0.323		0.224

16.2.4.4

Loading and Hauling

Production will be carried out using a fleet of 220-tonne class capacity haul trucks and hydraulic shovels with a bucket capacity of 26 m³ in ore and 26 m³ to 29 m³ in 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 four (4) pass-loadings of trucks hauling ore and waste and five (5) to six (6) pass-loadings of trucks hauling overburden. A maximum of 22 haul trucks will be required in the peak years.

The maximum shovel productivity per shift has been estimated at 30,300 tonnes of ore, 35,900 tonnes of waste per shift and 22,500 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 production equipment but also as a replacement for the shovel in down-time situations, as well as for other tasks involving material handling.

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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.

Table 16-12: Truck Speed and Fuel Consumption (Loaded and Empty)


	Haul Truck Loaded
Parameter	Acceleration 100 m		Flat (0%) Topo		Flat (0%) In-Pit/ Crusher/ Dump		Slope Up (10%)		Deceleration 100 m
Speed (km/h)		20		40		40		13		20
Fuel consumption (litres/hr)		393		150		200		375		27

	Haul Truck Empty
Parameter	Acceleration 100 m		Flat (0%) Topo		Flat (0%) In-Pit/ Crusher/ Dump		Slope Down (-10%)		Deceleration 100 m
Speed (km/h)		25		50		50		25		25
Fuel consumption (litres/hr)		118		118		118		27		27

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The calculated cycle times were based on round-trip haulage profiles, haul truck speeds, and load/stop dump times determined for each material. A trend of cycle time for each material type over the LOM is shown in Figure 16-18.

LOGO

Figure 16-18: Cycle Time Trend over LOM

16.2.4.5

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 (2020 - 2022), primary equipment requirements will be as follows:

•

22 x 220-tonne class diesel haul trucks;

•

2 x 26m³ diesel-hydraulic shovels;

•

1 x 29m³ electric-hydraulic shovel;

•

1 x 18m³ front end loader; and

•

3 x 8 ¹/₂” DTH blast hole drills.

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The haul truck fleet is shown in Figure 16-19 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 reclamation work.

LOGO

Figure 16-19: 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 New Gold based on experience in similar open pit mining operations. The requirements for 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 primary equipment replacement is required. After 2025 (Year 9), the final pit is depleted and most pieces of equipment are no longer needed. Hence, during the stockpile re-handling period, less equipment is required and utilization is reduced significantly. The only equipment replacement will be the used auxiliary loader in 2020 (Year 4) that will be used mainly in the aggregate plant area.

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Table 16-13 shows the annual mine equipment fleet requirements over the open pit production stage to support the mining operation for each year.

During the open pit post-production stage (stockpile reclaim from 2025 to 2030), fewer equipment will be required to support activities related to stockpile re-handling, site rehabilitation and site maintenance. The necessary post-open pit equipment is: three (3) trucks (two (2) operating and one (1) back-up), one (1) diesel-hydraulic shovel, one (1) large wheel loader as a replacement for the shovel in down time, three (3) track-dozers, one (1) wheel-dozer and one (1) motor-grader. The mine will also keep some key equipment that will be operated only on an as needed basis, such as one (1) water truck and one (1) fuel truck.

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Table 16-13: Annual Open Pit Mine Equipment Requirements


	Pre-Production Period				Production Period																		Stockpile Rehandling Period
	2015		2016		2017		2018		2019		2020		2021		2022		2023		2024		2025- Q1 Q2		2025- Q3 Q4		2026		2027		2028		2029		2030
Haul Truck Fleet
Haul Truck		6		7		13		19		21		22		22		22		20		12		8		2		2		2		2		2		2
Shovel Fleet
Hydraulic Shovel (Diesel)		2		2		2		2		2		2		2		2		2		2		2		1		1		1		1		1		1
Hydraulic Shovel (Electric)						1		1		1		1		1		1		1		1
Drill Fleet
Blastholes Drill		1		2		3		3		3		3		3		3		2		2		1
DTH Drill (Reverse Circulation Sample, Pre- Split)		2		2		2		2		2		2		2		2		2		1		1
Support Fleet
Wheel Loader		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1
Rental Loader										1		1
Motor Grader		1		2		3		3		3		3		3		3		3		3		2		1		1		1		1		1		1
Track-Dozer 580 HP		3		4		5		5		5		5		5		5		5		5		2		2		2		2		2		2		2
Track-Dozer 580 HP / for dyke construction and site rehabilitation		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1
Wheel-Dozer 680 HP						1		1		1		1		1		1		1		1		1		1		1		1		1		1		1
Auxiliary Fleet
Compactor						1		1		1		1		1		1		1		1		1		1		1
Water Truck (30 KL)		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1
Water Truck (75 KL)		1		1		1		1		1		1		1		1		1		1		1
Fuel/Lube Truck		2		2		2		2		2		2		2		2		2		2		1		1		1		1		1		1		1
Boom Truck		1		1		2		2		2		2		2		2		2		1		1

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	Pre-Production Period				Production Period																		Stockpile Rehandling Period
	2015		2016		2017		2018		2019		2020		2021		2022		2023		2024		2025- Q1 Q2		2025- Q3 Q4		2026		2027		2028		2029		2030
Wheel Loader (Used)		1		1		1		1		1		1		1		1		1		1		1
Tow Haul Truck (Used)		1		1		1		1		1		1		1		1		1		1		1
Hydraulic Crane, truck-mounted		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1		1
Dewatering Pump (100 HP electric)		1		2		2		2		2		2		2		2		2		2		2
Mobile Pump (150 HP)		2		2		4		4		4		4		4		4		4		4		4
Service Truck		1		1		1		1		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		1		1		1		1
Mini Bus		1		1		2		2		2		2		2		2		2		1		1
Pick-up Truck Crew Cab		6		12		12		12		12		12		12		12		12		6		6		3		3		3		3		3		3
Stemming Loader		1		1		1		1		1		2		2		2		2		1		1
Cable Reeler (Used)						1		1		1		1		1		1		1		1		1
Aggregate Plant		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		2		2		2		2		2		2
Total

Total Primary Fleet		9		11		19		25		27		28		28		28		25		17		11		3		3		3		3		3		3

Auxiliary Equipment		34		43		54		54		55		56		55		55		55		45		40		17		17		16		16		15		15

Total Mining Equipment		43		54		73		79		82		84		83		83		80		62		51		20		20		19		19		18		18

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16.2.5

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 267 in 2021 (Year 5). 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 amount of equipment to maintain. The ratio of maintenance/operation for the hourly employees is approximately 0.6 during the normal years of operation. This ratio assumes that maintenance activities such as rebuilding components and machining will be performed off-site. The hourly personnel calculation also includes a 5-weeks allocation for VSA (vacation, sickness and absenteeism).

Post-operation activities (from Years 2026 to 2030), 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 27.

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Table 16-14: Annual Hourly Personnel Requirements


	Pre-Production Period						Production Period																Stockpile Rehandling Period
	2015		2016		2017		2018		2019		2020		2021		2022		2023		2024		2025- Q1 Q2		2025- Q3 Q4		2026		2027		2028		2029		2030
Operations
Shovel Operators		6		6		12		12		12		12		12		10		8		4		4		4		4		4		4		4		4
Loader Operators		2		2		2		2		8		8		6		2		2		2		0		0		0		0		0		0		0
Haul Truck Operators		17		21		46		61		69		70		75		74		54		30		22		6		6		6		6		6		6
Drill Operators		9		10		15		16		17		17		17		17		14		8		7		0		0		0		0		0		0
Dozer Operators		15		15		23		23		23		23		23		23		23		23		12		4		4		4		4		4		4
Grader Operators		4		4		12		12		12		12		12		12		12		10		4		2		2		2		2		2		2
Water Truck Operator/ Snow Plow/ Sanding		4		4		4		4		4		4		4		4		4		4		4		0		0		0		0		0		0
Auxiliary Equipment Operators		6		8		10		10		10		10		10		10		10		10		6		2		2		2		2		2		2
General Labour		4		4		6		6		6		6		6		6		6		6		4		0		0		0		0		0		0
Janitor		2		2		2		2		2		2		2		2		2		2		1		1		1		1		1		1		1

Hourly Open Pit Operations Total		69		76		132		148		163		164		167		160		135		99		64		19		19		19		19		19		19

Field Maintenance
Field General Mechanics		4		6		8		12		14		16		16		16		14		10		4		0		0		0		0		0		0
Field Welder		2		2		6		6		8		8		8		8		8		6		4		0		0		0		0		0		0
Field Electrician		4		4		6		6		8		8		8		8		8		6		4		0		0		0		0		0		0
Shovel Mechanics		4		4		8		12		12		12		12		12		12		6		4		2		2		2		2		2		2
Shop Maintenance
Shop Electrician		4		4		5		5		5		5		5		5		4		4		2		0		0		0		0		0		0
Shop Mechanic		6		6		12		18		20		22		22		22		16		8		6		2		2		2		2		2		2
Mechanic Helper		2		2		10		10		10		10		10		10		6		3		1		1		1		1		1		1		1
Welder-machinist		2		2		4		6		6		6		6		6		4		4		1		0		0		0		0		0		0
Lube/Service Truck		4		4		5		5		5		5		5		5		4		4		2		0		0		0		0		0		0
Electronics Technician		2		2		2		2		2		2		2		2		2		2		1		0		0		0		0		0		0
Tool Crib Attendant		2		4		4		4		4		4		4		4		4		2		2		0		0		0		0		0		0
Janitor		2		2		2		2		2		2		2		2		2		2		2		0		0		0		0		0		0

Hourly Mine Maintenance Total		38		42		72		88		96		100		100		100		84		57		33		5		5		5		5		5		5

Hourly Personnel Total		107		118		204		236		259		264		267		260		219		156		97		24		24		24		24		24		24

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Table 16-15: Salaried Open Pit Personnel Requirements


	Pre-Production Period				Production Period																		Stockpile Rehandling Period
	2015		2016		2017		2018		2019		2020		2021		2022		2023		2024		2025 Q1 Q2		2025- Q3 Q4		2026		2027		2028		2029		2030
Operations
Open Pit Superintendent						1		1		1		1		1		1		1		1		1
General Mine Foreman		1		1		2		2		2		2		2		2		2		2		1
Mine Shift Foreman		4		4		8		8		8		8		8		8		8		4		4
Drill & Blast Foreman				1		1		1		1		1		1		1		1		1		1
Blaster		1		2		2		2		2		2		2		2		2		1		1
Blaster Helper		1		2		2		2		2		2		2		2		2		1		1
Dispatcher		2		4		4		4		4		4		4		4		4		2		2
Training Foreman		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		10		17		23		23		23		23		23		23		23		15		13		0		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		1
Maintenance Planner				2		2		2		2		2		2		2		2		2		1
Mechanical Engineer				1		1		1		1		1		1		1		1		1		1
Maintenance Foreman		1		2		4		4		4		4		4		4		4		4		2		1		1		1		1		1		1
Maintenance Trainer		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		3		8		11		11		11		11		11		11		11		11		7		1		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		1
Open Pit Engineer		1		1		1		1		1		1		1		1		1		1		1
Geotechnical Engineer		1		1		1		1		1		1		1		1		1		1
Blasting Engineer				1		1		1		1		1		1		1		1		1
Mining Engineering Technician		1		1		2		2		2		2		2		2		2		2		1
Mine Surveyor		1		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		1

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	Pre-Production Period						Production Period																Stockpile Rehandling Period
	2015		2016		2017		2018		2019		2020		2021		2022		2023		2024		2025 Q1 Q2		2025- Q3 Q4		2026		2027		2028		2029		2030
Salaried Mine Engineering Total		6		9		11		11		11		11		11		11		11		11		6		1		1		1		1		1		1

Geology
Chief Geologist						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
Grade Control Geologist		1		1		2		2		2		2		2		2		2		2		1
Geology Technician		1		1		2		2		2		2		2		2		2		2		1

Salaried Geology Total		2		3		6		6		6		6		6		6		6		6		3		1		1		1		1		1		1

Total Salaried Staff		21		37		51		51		51		51		51		51		51		43		29		3		3		3		3		3		3

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16.3

Underground Mining Methods

The proposed underground mine design supports the extraction of 1,500 t/d of ore by longitudinal longhole open stoping (“LHOS”), a mining technique suitable for the geometry and ground conditions of the Rainy River underground resources. Backfilling with cemented aggregate fill (“CAF”) is a significant aspect of the project with respect to maximization of both resource recovery and mining productivity. Modern trackless equipment will be employed in the majority of mining activities.

A main decline from a surface portal located to the east of the open pit will be used to access the mine. A fleet of 7 m³ Load Haul Dump trucks (“LHDs”) and 45-tonne trucks will be used for material loading and transport from the various underground working areas through an internal ramp system that connects all levels to the main decline. Loading will occur in close proximity to the stoping areas and ore will be hauled directly to a surface coarse ore stockpile adjacent to the portal.

An extensive underground development program that attains a peak of 560 m/month of jumbo advance is required to develop and maintain access to adequate resources to sustain 1,500 tpd of ore production. The ramp-up period to full production will require five (5) years from the onset of mine development. Waste generated through infrastructure development will be disposed of in underground stopes whenever operationally practical, however, an estimated excess of 1.80 Mt must be hauled to the surface waste stockpile over the life of mine.

Key mine infrastructure includes a backfill delivery raise that terminates at an underground truck loading station, a cement storage and grout mixing facility, two (2) main dewatering stations, an equipment maintenance facility, electrical substations, and other smaller, ancillary installations. Permanent fans located on surface will provide fresh air to the mine. A propane air heating system will be used during the winter months.

Ore occurs in subvertical horizons in varying widths from about 3 m to 20 m. Widths over 15 m are rare, and the weighted average across all zones is approximately 8 m. The ore footwall (“fw”) and hangingwall (“hw”) generally dip at 60 degrees or more, but can flatten locally to as low as 45 degrees in some areas.

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Rock mass conditions are projected to be generally very good and little water ingress is anticipated.

16.3.1

Mine Design

16.3.1.1

Access and Mine Infrastructure

A 4 km main decline driven at a -15% gradient from the surface portal will provide access for personnel and equipment. The main decline will connect to independent internal production ramps that will service all levels in each mining area. In the interest of good visibility, minimizing haulage time and minimizing equipment wear, the length of straight segments was maximized in the design of the decline.

Internal production ramps are of spiral configuration, and will be generally driven at a -10% gradient to provide level access on each revolution. A 25 m minimum turning radius was employed. Passing bays, remucks and safety bays were incorporated into the decline and all internal production ramps.

The mine design includes a significant amount of raise development to establish and extend the primary air circuit. Three (3) raises to surface (two (2) exhaust raises and one (1) fresh air raise) will be required. As the mining levels are developed from the internal ramps, raising from level to level will be necessary to permit the exhaust of contaminated air by advancing the primary circuit in increments. A backfill raise from surface has also been included for the delivery of aggregate to an underground CAF plant.

Mine infrastructure will also include:

•

A workshop for the maintenance and repair of underground equipment

•

An explosives magazine and a cap magazine

•

A permanent refuge/lunchroom

•

Two (2) refueling locations for mobile equipment

•

A main substation near the workshop area

•

Two (2) main dewatering stations and a suite of local settling sumps

•

Definition drilling drives and cubbies where required

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16.3.1.2

Level Development

Sublevels will be accessed from the ramps on a 20 m vertical interval that is defined by the planned stoping heights. Ramp development will be set back a minimum of 30 m from the ore contact. This arrangement promotes long-term geotechnical stability and provides adequate space for the placement of a return air raise and other services such as sumps, remucks and portable refuge locations.

Remucks on all levels will be used for the placement of blasted ore during the mining cycle. During the backfilling cycle, CAF and waste rock will be tipped here for rehandling into the stopes by LHD.

Ore drives will be developed along strike in two (2) directions from a common access point to the strike extent of the zone. In several areas, where multiple ore horizons are present, these ore drives will cross a waste divide to access stopes in the other horizons. The arrangement minimizes waste development as traditional fw drives along the strike of the orebody are not required. However, in the absence of fw drives, definition drilling must be accomplished from waste drives dedicated to this activity. The savings in waste development compared to an fw drive arrangement is significant.

Ore drives are designed to follow mineralized horizons even when they are not within ore-grade areas where stoping has been planned. This approach will generate approximately 900 kt of low-grade material (below the 3.5 g/t Au eq mining cut-off grade, but above 1.5 g/t Au eq); this material will be processed given that the mining costs are sunk and a profit can still be made from a marginal cost perspective. Material generated from the ore drives below 1.5 g/t Au eq will be retained u/g as backfill.

Figure 16-20 shows a sublevel arrangement in cross-section and Figure 16-21 shows typical level development requirements.

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LOGO

Figure 16-20: Sublevel Arrangement in Cross-section

LOGO

Figure 16-21: Typical Level Plan

Development design standards considered equipment size, services, and required activity. Ontario mining regulations require that main haulageways regularly used by pedestrians be at least 1.5 m wider than the maximum width of a motor vehicle. Furthermore, if the haulageway is

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less than 2 m wider than the maximum width of the vehicle, safety bays must be included. The widest mobile equipment at New Gold, the 45-tonne truck, is 3.18 m in width. Therefore, haulageways (designed at 5 m width) with truck and pedestrian traffic such as the main and production ramps, include safety bays.

To provide adequate overhead clearance between equipment and services such as ventilation ducting, a 5 m high drive is required. The main ramp, internal production ramps, level accesses and most ventilation drives are 5 m high by 5 m wide.

Ore drives were also designed at 5 m high by 5 m wide. This width will permit adequate spacing for a remote stand for remote stope loading. AMC recommends that cut-outs for remote stands should also be considered relative to what may be seen as common and best practice in Ontario mines. The height and width also provide sufficient operating clearance for the production drill rig.

Development design parameters are summarized in Table 16-16 and Table 16-17. Figure 16-22 and Figure 16-23 illustrate standard designs for ore drives and the main decline, internal production ramps and level accesses, respectively.

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Table 16-16: Lateral Development Design Parameters


	Lateral
Parameter	Remuck		Ore Drive		Level Access		Main Decline		Expanded Main Decline		Exhaust Drift		Production Ramp		Passing Bay		Definition Drilling Drift		Safety Bay		Substation		Sump
Width (m)		5		5		5		5		6		5		5		4		5		1.5		5		5
Height (m)		7		5		5		5		6		5		5		5		5		2		5		4
Length (m)		15		—		—		—		—		—		—		20		—		1.8		8		8
Max Gradient (%)		2		2		2		15		15		15		15		15		2		2		2		15

Table 16-17: Vertical Development Design Parameters


	Vertical
Parameter	Fresh Air Raise			Return Air Raise			Ventilation Drop Raise		Alimak Raise		Backfill Aggregate Raise
Width (m)		3	*		3	*		3		3		2	*
Height (m)		—			—			3		3		—
Length (m)		—			—			—		—		389

Diameter

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LOGO

Figure 16-22: Standard Design – Ore Drive

LOGO

Figure 16-23: Standard Design – Main Decline, Internal Ramps and Level Accesses

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16.3.2

Stope Design

AMC used the Mineable Shape Optimizer (“MSO”) module from the Datamine Studio 3 mine planning software package to produce conceptual stope shapes. Key design parameters used in MSO are summarized in Table 16-18. The conceptual stope shapes were refined as necessary to minimize the amount of planned dilution and to meet practical mining constraints.

Table 16-18: Stope Design Parameters


	Parameter	Units	All Zones
MSO Parameters	Au eq Cut-off*	gpt		3.5
	Level Spacing	m		20
	Stope length increments	m		2
	Minimum Mining Width	m		3
	Minimum Waste Pillar Width	m		10
	Minimum Footwall Dip	degrees		55
	Minimum Hanging Wall Dip	degrees		55
	Footwall Overbreak	m		0.25
	Hangingwall Overbreak	m		0.25
Au eq Parameters	Au (Gold) price	$/oz		1,250
	Au (Gold) recovery	%		88
	Ag (Silver) price	$/oz		20
	Ag (Silver) recovery	%		75

Isolated areas meeting cut-off grade were evaluated against access development costs to determine economic viability, before including them in the Mineral Reserves.

The Intrepid lower, 17 East lower, ODM West, and the upper portion of the 433 Zone contained economically marginal material that required individual analysis to determine profitability. Considering the capital development required to access and mine this material, the analysis indicated that this marginal material carried sufficient profit margin and was therefore kept within the current mine plan.

The LOM plan includes approximately 520 stopes across all mining areas. Figure 16-24 and Figure 16-25 are long-section views showing stope shapes generated by the MSO process.

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LOGO

Figure 16-24: Mineable Stope Shape – Intrepid Zone

LOGO

Figure 16-25: Mineable Stope Shape – ODM and 17 East Zones

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16.3.3

Mining Method and Sequence

16.3.3.1

Zone Definition

The orebody zones were defined by spatial location and elevation, which facilitate 1,500 t/d of production through the creation of multiple working areas. The ore footwall (“fw”) and hanging wall (“hw”) generally dip at 60 degrees or more, but can flatten locally to as low as 45 degrees in some areas. The maximum lateral extent of the ore bodies varies significantly from 570 meters, in the ODM Main Zone to 75 m in the 17 East Upper Zone. Figure 16-26 shows the zone locations.

The Intrepid Zone contains an area of waste material effectively creating a natural sill that further divides the area into an upper and lower zone. To allow production in the ODM Main Zone without having to develop to the bottom of the zone first, a sill pillar, located in lower grade ore, is included, creating an upper and lower ODM Zone.

LOGO

Figure 16-26: Mining Zones

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16.3.3.2

Stope Cycle and Sequence

The mining method will be longitudinal LHOS, retreating from the strike extent of ore on each level. Stopes are typically 20 m in length along strike and 20 m high. Stope widths (thickness of mineralization) range from about 3 m to 20 m but average approximately 8 m.

From each level access, drives will be developed within the mineralized horizon along strike on the top sill and bottom sill levels. Force ventilation will provide fresh air to the working faces. Stope production begins with a drop raise followed by a slot blast and production ring drilling. Production blasting and mucking will proceed cyclically until the stope is depleted and all ore has been mucked out.

Longitudinal LHOS is a non-entry method, with remote mucking of blasted ore required once the drawpoint brow is open to the extent where the operator may be exposed to uncontrolled sloughing from the stope cavity. To protect the operator from the remotely operated loader, a remote mucking stand is the minimum requirement; an excavated cubby may also be desirable.

The remote mucking stand is a portable prefabricated platform designed to withstand the impact of a loader in the event that control has been lost by the operator. Past practice has generally involved securing the stand to the drift sidewall for stability. Increasingly, cubbies are excavated into the drift sidewall to position the operator off any potential line of travel of the loader.

AMC recognizes that both approaches are currently used in Ontario and does not prescribe which approach should be adopted by New Gold. AMC has used the prefabricated remote mucking stand in this study for costing purposes only. Ultimately, remote mucking procedures must be established by New Gold through risk assessment and consideration of (best) safe work practices.

Once mucking of blasted ore is complete, backfilling commences with the placement of a sufficient volume of CRF to provide stability of the backfill that is re-exposed during the extraction of the next stope in sequence. Once this volume of CRF has been placed, uncemented waste from underground development or from surface is used to fill the remaining void – see Figure 16-27 and Figure 16-28.

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LOGO

Figure 16-27: Typical Stope Long Section

LOGO

Figure 16-28: Typical Stope Cross-sections

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Mining proceeds upwards from the lowest level of the zone or from an adopted sill elevation, with backfill providing the working platform for each successive lift. To achieve and maintain 1,500 t/d of production, multiple zones will be concurrently active.

16.3.3.3

Backfilling Options

An assessment of backfilling options was completed as part of the Feasibility Update Study. The proposed mining method, filling cycle time, availability of fill materials, capital cost and operating costs were the main criteria in the selection of the backfill option. The backfill types investigated include aggregate fill (“AF”) and cemented aggregate fill (“CAF”), rock fill and cemented rock fill (“RF & CRF”), paste fill (“PF”), hydraulic fill and cemented hydraulic fill (“HF/CHF”). AMC has estimated cement-dosing rates based on experience with similar operations. The outcome of the options study indicated the following conclusions and recommendations.

Conclusions:

•

There are a limited number of mining areas available at any time such that a short mining cycle time is essential to achieve production targets. The CAF system has the shortest filling cycle time;

•

The CAF system has relatively low up-front capital and acceptable operating costs. The crusher dedicated to surface road aggregate and stemming production will also produce aggregate for underground backfill; and

•

Implementation time for the CAF fill system is minimal, requiring only a few months to complete the surface batch plant, the underground grout delivery reticulation and the underground loading and spraying system.

Recommendations:

•

A CAF system should be implemented for backfilling the underground mine;

•

CAF trials should be conducted to enable an initial assessment of the quality of the aggregate that can be produced from crushed waste rock;

•

The trials should target the preparation of CAF samples containing 3% w/w to 6% w/w cement using a cement slurry of approximately 55% w/w cement, or a water to cement ratio of 0.8;

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•

These trial samples should be sealed and cured in a humid environment and strength-tested after 3, 5, 7, 14, 28 and 56 days; and

•

Confirm bulk density of the aggregate over a number of samples.

16.3.3.4

Proposed CAF System Description

Backfilling with cemented aggregate is an integral part of the mining cycle. The void to backfill was based on the average 1,500 tpd planned ore production. This corresponds to an average daily void to backfill of approximately 550 m³. A stockpile of crushed and graded aggregate on surface will feed an aggregate fill pass down to 400 metres below surface. The fill pass, 2.0 m in diameter, will be constantly fed and maintained full at all times.

The surface infrastructure consists of a 100 tonne cement silo, a calibrated transfer system and a small 4 m³ agitated slurry tank to prepare batches of cement slurry. On demand from underground, the slurry will be discharged to the underground holding tank via a 2” slick line in a borehole. After each batch, the line will be flushed automatically, with the flush water being diverted to the sump in the underground backfill station.

The underground infrastructure consists of an aggregate loading station and cement slurry system. A loading chute fitted to the base of the aggregate raise will be used to fill the backfill truck. As the driver operates the chute, a batch of cement slurry will be sprayed onto the aggregate as it falls into the truck, pumped from a 6 m3 agitated slurry tank. The volume of cement slurry will vary with the required cement dosage for that batch.

The driver will then transport the CAF to a mixing bay adjacent to the entry cross cut on the level where fill is required and tip the load into the mixing bay.

An LHD will remix the CAF in the mixing bay and then pick up a bucket load of CAF and transport it to the tipping point at the stope. The LHD will approach the stop-block slowly, then tip a bucket load of CAF into the stope. This process will be repeated until the CAF rill reaches the floor elevation of the drift or the required volume of CAF has been placed.

If additional CAF is required and it is impractical to tip it over the stop-block, then the fill operation will change to remote loading operations to ensure the safety of the operator. The stop-block will be removed and, using remote loading procedures, the next two (2) CAF buckets will be tipped on the floor in front of the rill. The LHD will then push the two (2) piles forward over the rill, taking care not to overextend. This operation will continue until the required volume of CAF has been placed.

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The remaining stope volume can then be filled using the same remote loading procedure with uncemented development waste rock until the fill has reached the level of the fill horizon. In some high grade stopes, a thin layer (~0.5 m) of CAF may be placed on top of the RF to achieve a distinct mucking surface and reduce loss of ore into the RF. Figure 16-29 shows a schematic of the CAF system.

LOGO

Figure 16-29: CAF System Schematic

16.3.4

Waste Management and Stope Filling

Considerable waste rock will need to be disposed of on an ongoing basis throughout the mine life. Stopes will be filled with uncemented development waste rock wherever possible and as required, but some waste will inevitably be hauled to surface for disposal in waste stockpiles. As a priority, waste rock from development faces will be placed as required in available stopes voids. If voids are unavailable, waste can be stockpiled underground, as capacity permits.

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As development activity will slow down in the last two (2) years of production, sufficient development waste will not be generated to provide material for backfilling. The balance will be made up with uncemented crushed aggregate. Table 16-19 tabulates the mass of waste to be generated from development headings and the destination of these volumes over time. Over the LOM, 39% of development waste ore is estimated to be placed back underground. The balance will be disposed of in surface waste stockpiles.

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Table 16-19: Life of Mine Backfilling Waste Rock


Year	Ore Tonnes (‘000 t)		Waste Tonnes Developed (‘000 t)		Waste Tonnes as RF (‘000 t)		Waste Tonnes to Surface (‘000 t)		Crushed Aggregate Required for RF (‘000 t)		CAF Tonnes Required (‘000 t)		Cement Tonnes Required (‘000t)
2013		0		0		0		0		0		0		0
2014		0		0		0		0		0		0		0
2015		0		106		0		106		0		0		0
2016		16		382		0		382		0		0		0
2017		198		369		40		329		0		48		2
2018		237		429		77		353		0		92		4
2019		425		319		117		202		0		140		6
2020		551		231		147		85		0		176		7
2021		552		244		159		85		0		191		8
2022		554		290		193		97		0		232		9
2023		550		301		199		101		0		239		10
2024		543		231		173		59		0		207		8
2025		443		16		16		0		177		231		9
2026		119		0		0		0		47		56		2

Total		4,187		2,920		1,121		1,799		224		1,613		65

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16.3.5

Development and Production Schedule

16.3.5.1

Production Rate

AMC completed a stope cycle time analysis on parameters representative of the average stope size to define the maximum schedulable rate that any given stope may be extracted in the LOM plan. The results of this exercise yielded the following:

•

Stopes are mined at 350 tpd once development is complete.

•

Stopes are backfilled at 500 m³/day once mining is complete.

In each zone, the order of stope extraction was then linked in a sequence that respects the constraints of the mining method and the development that must be accomplished before any given level may be put into production. A preliminary schedule was created based on the maximum schedulable rates.

The preliminary schedule was then refined through an iterative process to:

•

Maintain a constant 1,500 tpd production rate following the ramp up period;

•

Optimize the grade profile over the LOM, favouring higher grade areas earlier; and

•

Accomplish the above with levelled and reasonable development rates and resources.

AMC concludes that 1,500 tpd is sustainable, and that the target rate is a reasonable estimation of the capacity of the underground mine in consideration of development requirements.

16.3.5.2

Pre-Production Development

Pre-production development will span a 24-month time frame before the first ore is mined in the Intrepid Zone. The ramp-up period to full production will require five (5) years from the onset of mine development.

Given its proximity to surface, the development strategy targets the Intrepid upper block as the first priority, followed by the more distant zones to achieve full production. Hence, the first stope will be extracted from the Intrepid upper block.

Critical path pre-production activities include:

•

Portal construction and development of the main decline to the Intrepid Zone.

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•

Establishment of the Intrepid ventilation raise to support the advancement of the main decline towards the ODM zone and provide secondary egress from the Intrepid area.

•

Partial development of the Intrepid production ramp and levels.

•

Definition drilling of the Intrepid Zone

The Intrepid Zone will be the unique source of ore production in 2019, and will still represent the largest (75%) source of production in 2020. It is not until 2021, the fifth year of underground development, that significant tonnage can be extracted from the ODM Zone.

Figure 16-30 illustrates the extent of development at the onset of stope production in the Intrepid Zone in April 2019. A total development requirement of approximately 7,200 lateral metres and 485 vertical metres is planned in the first 24 months. Up to 560 m/month of development advance will be required at the peak activity level. Relative to company strategy and goals beyond the feasibility study, AMC believes that opportunities may exist for further optimization of the development and production scheduling.

LOGO

Figure 16-30: Extent of Mine Development at the Onset of Production (April 2019)

16.3.5.3

Sustaining Development

The Upper Intrepid Zone will be mined to bring as much production into the earlier years as possible. However, the production rate achievable in the Intrepid Zone is constrained by a limited number of working areas such that achieving and sustaining full 1,500 tpd production is dependent on the development of the ODM zone. The ODM Main zone has been prioritized in the schedule given its size (Reserve tonnage) and the above-average grade.

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Key infrastructure that is required to develop the ODM Main Zone includes the westerly extension of the main decline by approximately 2 km, the main air intake raise, the main exhaust raise and the exhaust ramp to complete the primary circuit. The workshop area and CAF plant will also be completed during this phase.

The ODM Main Zone is divided into a lower mining block and an upper mining block, separated by a sill, to reduce development requirements before the zone can be put into production. The upper block has been prioritized as it requires the least development and contains the highest grade Reserves.

Two (2) other zones are accessed in parallel to the OMD Main zone, the 433 Zone and the ODM West zone. The upper area of the 433 Zone is under the final pit and will be mined as a priority before the pit has developed down to this level. The lower portion of the 433 zone will be mined as resources become available. The ODM West Zone is a smaller, yet high grade zone located in the western extent of the mine.

Significant development is required to access and ventilate the 17 East Lower Zone. Its low grade dictates a lower priority and is developed as resources become available.

The Lower Intrepid Zone is low grade and narrow and will be delayed until toward the end of the mine life. The last zone to be mined is the 17 East Upper. This will be accessed with an adit into the pit wall and will be delayed to avoid concurrent development with the open pit.

16.3.5.4

LOM Production Schedule

Full 1,500 t/d production will be effectively achieved in 2022.

Figure 16-31 illustrates the ramp-up to full production tonnage as the various zones are brought into production.

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LOGO

Figure 16-31: Underground Production Profile

Figure 16-32 shows the LOM split of production by development and stoping.

LOGO

Figure 16-32: Production by Activity

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16.3.6

Geotechnical

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, 2014A) and recommended to Australian Mining Consultants (AMC) to support underground mining engineering. Figure 16-33 provides a North West view of the underground mine geometry developed for stress modelling.

During the 2012 drilling campaign (AMEC, 2013E), three (3) 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. As part of this Feasibility Study, New Gold additionally performed orientation of cores for four (4) boreholes in the Intrepid Zone. These holes, including other selected exploration cores, were subsequently geomechanically logged by AMEC to provide additional supporting data for geomechanical design (AMEC, 2014A).

LOGO

Figure 16-33: View North West of the Underground Mine Geometry

Developed for the Map3D Numerical Stress Modelling, Indicating Zones and

Main Underground Boreholes (AMEC, 2013G).

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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, were obtained characterizing the hangwall (“hw”), ore zone (“oz”) and footwall (“fw”) domains of the largest west zone respectively. Typical rock mass qualities in the Intrepid Zone had average Q’ values of 21, 22 and 17, were obtained for the hw, oz and fw domains. 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 of the ODM based on the present data. Additional, above and to the east of the Intrepid Zone, there is a zone of brecciated rock that is found to be developed in sub horizontal structures that terminate rapidly. These also have a lower RQD in the range of 10 to 70 (average of 40), and an average Q’ of ~ 4, however, mining does not intersect these zones and the crown pillar thickness is upward of greater than 50 m and considered to be stable.

Intact failure curves were developed based on laboratory testing of 30 UCS, 27 triaxial tests and 24 Brazilian tensile tests. The overall average UCS results, used for linear elastic numerical stress analysis using the Hoek and 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 and 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 (Potvin, 1988; Hadjigeorgiou et. al., 1995), Longitudinal LHOS retreat stope panels will be 20 m along strike in all zones and the LHOS shallower than 500 m will not require cable bolt support to maintain back or hangingwall stability at shallow depths (Figure 16-34). 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 stopes, however, this accounts for only 3 % of these deep stopes (AMEC, 2013G; AMEC, 2014A).

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LOGO

Figure 16-34: Modified Stability Graph for ODM West Zone above 500 m depth Indicating Stability

of Stope Surfaces Based on Potential Design Limits and Actual Final Stope Dimensions

[Longitudinal LHOS Retreat 20 mH x 20 mL x 12 mW average dip 59 degrees] (AMEC, 2013G)

Ground support for the underground mine development at the RRU will consist of three (3) packages of increasing complexity (Table 16-20) 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 some longhole stopes where increased levels of ground support have been assumed based on linear elastic modelling of the mine sequence. Overall, it is anticipated that approximately 5% of ore development above 500 m in depth will require shotcrete, while 15% of lateral ore development below 500 m depth will require shotcrete. Additionally, based on the numerical stress modelling, approximately 5% of the lateral ore development below 500 m, will require a high stress support system, which is recommended to consist of mechanized cone bolts (“MCB”), installed in a similar pattern to that

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in the Sudbury basin and Brunswick Mine (Simser et. al., 2002). Additionally, shotcrete is estimated to be required for 5% of all access development and 10% for the main ramp. Cable bolts are recommended in intersections with spans greater than 10 m.

Table 16-20: 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 DL 3³ 2.3 m ³/₄“ MCB on 1 x 1 m pattern installed in the back and side walls connected with 0 gauge screen strapping

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	As per D2 plus based on DL3³ 2.3 m ³/₄“ MCB on 1 x 1 m pattern installed in the back and side walls connected with 0 gauge screen strapping

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

^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 stopes will be partially backfilled with cemented aggregate fill (“CAF”) at an average binder dosage of 5%, the cemented backfill being placed at an angle of repose until level with the top cut, the rest of the stope being filled with uncemented rock fill. Testing indicates that a 3% binder consisting of 10% cement (NPC) and 90% Slag has acceptable strength development at 28 days of 1 MPa and no degeneration in strength was noticed for testing at 180 days (AMEC, 2013B, 2013v).

Stope dilution has been based on the empirical estimation of wall slough after Pakalnis (2002) (Figure 16-35). Depending on the ore zones, stope hw dip and depth, 0.25 m to 0.5 m of Equivalent Linear Overbreak/Sloughs (“ELOS”) are estimated, for a sublevel spacing of 20 m, and 20 m strike length for the majority of stopes. CAF backfill dilution from stope walls have been estimated at 0.3 m based on bench marking to similar operations (AMEC, 2013G).

LOGO

Figure 16-35: 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).

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Based on the present mine plan, no crown pillars will be required below the open pit, since the stopes directly below the pit will be mined in advance, they are relatively narrow and will be backfilled with cemented rockfill. The 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. As previously noted, a crown pillar will exist for the new Intrepid Zone, that based on the present mine geometry the crown pillar thickness of competent rock to the bedrock surface, above mining is expected to be between 45 to 100 m thick. Based on a mining width of 20 m by 6 m, this crown is considered stable. A sill pillar will be developed in the ODM/17 zone between -150 and -170 m elevation. This sill will be moderately stressed, but will be mined towards the end of mine life, however, only account for a limited number of stopes (5 stopes), in which a reduced extraction ratio has been assumed.

As some stress induced damage may be anticipated a 32 channel microseismic system is recommended for the deeper region of stopes and the region below the pit. Additionally, standard displacement monitoring and cable bolt monitoring instrumentation will be used.

16.3.7

Drill and Blast

16.3.7.1

Bulk Explosive Product

Ammonium nitrate and fuel oil (“ANFO”) is the most basic bulk explosive product. It is widely available and can be poured or blown into long holes. The equipment required to use ANFO is basic and easily maintained. ANFO is the primary explosive product that will be used for both the lateral development and stoping at the Rainy River mine.

16.3.7.2

Initiation System – Boosters and Detonators

Boosters are basic solid cartridge explosive products that are readily available from explosive products suppliers, and are planned for lateral development and stoping. Boosters can be initiated with high strength electric, electronic and non-electric detonators.

Electronic detonators are planned for blast initiation (i.e. starter cap) of lateral development and stopes. They have multiple security features that make them virtually immune to interference, unauthorized use, or unplanned detonation.

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16.3.7.3

Longitudinal Down-Hole Stope Design

Figure 16-36 and Figure 16-37 show a typical long section and cross section of the longitudinal down-hole stope of a height of 20 m (floor to floor), average width of 8 m (Widths range from 3.5 m to 15 m, but 8 m is examined), and length of 20 m. Longitudinal down-hole stopes account for the majority of the planned stope tonnage (approximately 97%). A minor amount of transverse stoping and uppers will be required but, given their very minor contribution to the LOM plan, the detail is not examined here.

LOGO

Figure 16-36: Longitudinal Downhole Stope Long Section

LOGO

Figure 16-37: Drop Raise Drillhole Pattern for the Longitudinal Downhole Stope (Section View)

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16.3.7.4

Explosive Consumption

The Rainy River underground mine will consume approximately 42 tonnes/month of ANFO, 4700 detonators/month, and 4700 boosters/month, which is based on the peak production rate of 1,500 tpd and a development advance rate of 8 m/d. A small reserve of cartridge explosive products will also be required for special situations (i.e. wet holes in long hole stopes and development rounds).

16.3.8

Mobile Equipment Requirements

16.3.8.1

Underground Contract Mining Phase

A contractor will be engaged to provide experienced personnel and expertise during the first three years of underground activity and will be responsible for all development and mining activities during this period. The underground mining equipment will be supplied by New Gold. Table 16-21 shows the equipment build-up during this period.

The transition to New Gold personnel underground will likely commence during 2019, the third year of development. However, for the purpose of cost estimation, AMC has assumed contract labour rates for all manpower in 2019.

Contractor assistance will extend beyond 2019 but will be limited to mechanized raise development (Alimak), the installation of ladderways and any residual construction to complete the workshop area.

Temporary maintenance capacity for mobile equipment will be required before commissioning of the permanent arrangement. During the contract mining phase, the temporary facilities will be provided for, and managed, by the contractor in the portal area. It may be advantageous to equip a second temporary workshop for minor and routine maintenance closer to the working areas until the permanent workshop is completed in 2020.

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Table 16-21: Mobile Equipment Requirements – Contract Mining Phase


Mobile Equipment	2017		2018		2019
Drill Jumbo, 2-Boom		1		2		0
Drill Longhole		1		0		1
Haulage Truck, 45 Tonne		2		0		1
LHD, 7.2m³with Remote		2		0		1
Bolter		1		1		0
Lubrication Service Truck		1		0		1
Boom Truck		1		0		0
Scissor Lift		1		0		0
Face Charger, Explosives		1		1		0
Pneumatic Cartridge Loader		0		1		0
Pneumatic ANFO Loader		0		1		0
Blasting Utility		1		0		0
Shotcrete Sprayer		1		0		0
Personnel Carrier		1		1		0
Transmixer		1		0		0
Motor Grader		0		1		0
Face Mapper		0		1		0
Fork Lift		0		1		0

Total		15		10		4

16.3.8.2

Production Phase – 2019 Onwards

During steady state operations, New Gold will supply the labour force and continue to supply all equipment with the exception of mechanized raise climbers “(“MRC”). MRCs will be used for internal ventilation raises in the ODM East and ODM Upper zones and are included in the raising contract. In addition to production and development equipment, support equipment is also required. Table 16-22 and Table 16-23 show the required equipment for development, stoping and support activities, respectively.

Utilizations were determined by calculating the number of hours per year required for each piece of equipment to achieve the targeted mine production and development rates. The hours were calculated using first principles. The utilization hours were then set against the total hours per

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year that the equipment was available. The available hours were determined by multiplying the estimated availability by effective working hours per year. To reduce the potential for production delays, additional unit(s) of key equipment that perform tasks directly related to production were, in some cases, scheduled to be purchased for redundancy.

Table 16-22: Underground Development and Production Equipment List


Description	Availability			Utilization						Qty
Description	Availability			Peak			Average			Qty
Two boom mining jumbo		85	%		50	%		50	%		3
LHD, 17 tonne (diesel - development/production/backfill)		85	%		80	%		80	%		4
Haulage Truck, 45 Tonne		85	%		80	%		80	%		6
Bolter		85	%		80	%		80	%		2
Top hammer longhole drill		85	%		50	%		50	%		2
Face Charger, Explosives Loader		85	%		80	%		80	%		2
Production Explosives Loader		85	%		50	%		50	%		1
Pneumatic Cartridge Loader, Explosives		85	%		50	%		10	%		1
Shotcrete Sprayer		85	%		10	%		10	%		1
Transmixer		85	%		30	%		30	%		1

Table 16-23: Support Equipment


Description	Availability			Utilization			Qty
Personnel Carrier, Diesel, Underground		85	%		30	%		3
Scissor Lift Truck, Diesel		85	%		30	%		1
Lubrication Service Truck, Diesel		85	%		30	%		2
Boom Truck, Diesel		85	%		35	%		1
Explosives Utility, (Transport)		85	%		30	%		1
Forklift		85	%		30	%		1
Utility Platform Lift		85	%		35	%		1

Jumbos

During steady state operations, an average of 475 m/month of lateral development must be achieved. A two boom jumbo capable of drilling holes up to 4.3 m deep was selected based on the average drift size.

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Based on first principles estimation of Jumbo productivities, three units are required.

Haulage Trucks and Loaders

A study was completed to determine which truck and loader combination would provide the highest productivity for the lowest total cost. Considering both capital and operating costs, the study concluded that 45 tonne trucks and 17 tonne loaders provide the lowest capital cost and operating cost (on a per tonne basis).

For the scheduled stope, development and CRF and RF volumes, four (4) 17-tonne loaders and six (6) 45-tonne trucks are required.

Bolters

Typical bolting patterns involve 2.4 m long resin grouted rebar on a 1.2 m square pattern on the back,1.5 m long resin grouted rebar on a 1.2 m square pattern in the walls and welded wire mesh. To maintain the targeted development rates, first principle productivities indicate two bolters are required. The bolter will be fitted with a mesh handling arm.

Longhole Drills

The primary function of the longhole drill is for production drilling. The longhole drill is also specified for cable bolt and drop raise drilling. First principle productivity estimates indicate that two longhole drills are sufficient to meet the projected demand.

Explosive Loaders

For development loading, two face charging units will be required to load as many as six (6) rounds per day. For production loading of up-holes and down-holes, a pneumatic ANFO loader is required. A pneumatic cartridge loader is also specified for the loading of wet upholes.

Shotcrete Sprayers

It is anticipated that approximately 5 - 15% of ore and waste development advance will require wet fibre-reinforced shotcrete. A minor amount will also be required for the construction of ventilation bulkheads and other infrequent applications. Although the anticipated volume is less than one cubic metre per day, the campaign nature of shotcrete spraying, the ability to meet development targets and consideration of operator safety supports the purchase of this unit, in contrast to the employment of hand sprayers. One (1) unit will meet the forecast volume.

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Transmixers

Shotcrete will be delivered from the surface batch plant to the underground headings in a transmixer. As shotcrete and concrete volumes are estimated to be low, only one (1) transmixer will be required.

Personnel Carriers

Personnel carriers will be needed for employee transport. The transporter selected is capable of carrying nine people. With three transporters, 27 people can be brought to their work places each shift. Personnel required underground on an as-needed basis, such as technical staff, are also transported underground in the personnel carriers.

Scissor Lift Truck

The scissor lift truck will be required to install and remove services (pipe reticulation, power cables and ventilation ducting) hanging fans, and assisting with construction. One scissor lift truck is specified. For redundancy, the fork lift fitted with a man basket and the utility platform lift are also capable performing similar tasks.

Lubrication Truck

A lubrication truck will be required to deliver grease, hydraulic oil and engine oil to equipment that is not likely to return to the shop area at frequent intervals. Utilization can be increased by keeping equipment near the working headings. This will reduce traffic on ramps and the main decline and reduce non-productive travel time for key equipment. This equipment would include trucks, LHDs, jumbos, longhole drills and, bolters. The service truck will travel between these equipment pieces to perform the required servicing.

Boom Truck

Bulk explosives (bagged ANFO) products will be transported from the surface to the underground powder magazine with a boom truck. This unit will also serve to transport materials from surface to underground. Material stockpiles will be set up throughout the mine for supplies such as rock bolts, screen, resin, pipes, vent duct, etc.

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Explosives Utility Vehicle

A Personnel Carrier will be dedicated to the transfer of explosives products from the surface magazines to the underground magazines. This includes packaged explosives products, detonating cord, non-electric caps, and electronic detonators.

Forklift

A fork lift with an integrated tool carrier will be required for general service activities. The integrated tool carrier can operate with a wide variety of accessories such as a man basket, forks and small loader bucket. Primary uses include, but are not limited to, handling of explosive pallets, providing personnel access to the backs and distributing road base material. One unit is specified.

Utility Platform Lift

The development cycle will include chip sampling and geological mapping of each round, to be accomplished safely in the absence of face bolting. The identification of bootlegs and face mark-up for drilling is also performed at this stage in the cycle. A utility platform lift has been dedicated to these activities, and it will be required to service approximately two rounds per shift on average, although variations in workload will vary.

16.3.9

Ventilation

The function of the ventilation system is to dilute/remove airborne dust, diesel emissions and explosive gases, and to maintain temperatures at levels necessary to ensure safe production throughout the life of the mine. The ventilation system has been designed to meet the requirement of the Occupational Health and Safety Act, R.R.O. 1990, Regulation 854 – Mines and Mining Plants (The Regulations).

The design is based on an exhausting system configuration with the main surface fans located at the Intrepid Exhaust and Primary Exhaust raises. Direct-fired mine air heating systems are located at the portal and Primary Intake raise. Fans are also located at these intakes to ensure balance of airflow across the propane burners.

Fresh air enters the mine through both the portal and primary fresh air raise with fresh air generally being distributed through the mine’s various ramp systems. Return air is exhausted from the mine through internal raises adjacent to each ore block. For the main production areas the return air progresses to a transfer drift and subsequently the Primary Exhaust Raise. Due to its distance from the other ore zones, the Intrepid Zone is exhausted independently to the surface.

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Level distribution is designed such that fresh air will be sourced from ramp accesses via auxiliary fan and duct into each operating level. Contaminated air from development and production level activities returns to an internal return air raise. A total of 335 m³/s is planned for ventilation of the Rainy River underground operation.

Two means of egress are provided for each production area of the mine. The primary means of egress is via the ramp system. Secondary emergency egress is provided in all internal raises and the Intrepid exhaust raise by means of installed ladderways. In the event of compromised egress to the surface through the main ramp, the Primary Intake Raise is fitted with an Alimak-Hek elevator. Figure 16-38 shows an isometric view of the Rainy River ventilation system.

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Figure 16-38: Rainy River Ventilation System

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16.3.9.1

Total Airflow Requirements

Air volume requirements are calculated to ensure safe production. The amount of air required is largely determined by the number and size of diesel equipment operating underground. The air volume supplied must be able to dilute and remove dust and noxious gases as well as diesel particulate matter generated by the use of such equipment.

The Regulations state that the ventilation quantity shall be at least 0.06 cubic metres per second for each kilowatt of power of the diesel-powered equipment operating in the workplace.

Primary infrastructure underground requiring continual ventilation is accounted for in the total airflow calculation. An airflow allowance is also determined for underground infrastructure and leakage and balancing inefficiencies.

Based on the scheduled production and development activities, airflow allocations are summarized in Table 16-24.

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Table 16-24: Total Airflow Requirements


Equipment	Number		kW		hp		Utilization¹			Total Airflow (m³/s)		Total Airflow (cfm)
Haul Truck		6		438		587		85	%		134		283,989
Scoop		4		321		430		85	%		65		138,753
Jumbo		3		110		147		15	%		3		6,293
Bolter		2		110		147		15	%		2		4,195
Longhole		3		110		147		10	%		2		4,195
Personnel Carrier		3		95		128		40	%		7		14,568
Scissor Lift		2		110		147		30	%		4		8,391
Lube Truck		1		120		161		40	%		3		6,102
Boom Truck		1		120		161		40	%		3		6,102
Blasters Truck		1		95		128		40	%		2		4,856
Shotcrete Sprayer		2		120		161		15	%		2		4,577
Face Charger		2		110		147		15	%		2		4,195
Transmixer		1		155		208		40	%		4		7,882
Grader		1		129		173		50	%		4		8,200
Face Mapper		1		96		129		15	%		1		1,831
Fork Lift		1		54		72		40	%		1		2,746
Sub-Total – Diesel Equipment											239		506,876


Infrastructure	Number		Unit Airflow (m³/s)		Total Airflow (m³/s)		Total Airflow (cfm)
Shop Complex		1		26		26		55,091
Fuel Bays		1		26		26		55,091
Magazines		0		0		0		0
Sub-Total – Infrastructure						52		110,182


Final Volume	Total Airflow (m³/s)		Total Airflow (cfm)
Sub-Total – Diesel & Infrastructure		291		617,058
Contingency/Leakage – 15%		44		92,559
Final Volume		335		709,616

¹	The utilization factor presented in this table reflects the expected peak ventilation requirements during a typical shift underground.

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16.3.9.2

Ventilation Modelling

AMC developed a ventilation model (using Ventsim) for the Rainy River Project for three (3) primary purposes:

•

To validate the operability of the ventilation circuit to ensure airflow can be provided to all the required areas.

•

To ensure compliance with design criteria.

•

To determine fan duties and energy requirements.

16.3.9.3

Permanent Primary Fans – Exhaust

Over the life of mine, the ventilation circuit will change depending on the type of activities and their location throughout the mine. AMC has modelled the circuit to reflect the peak primary fan duties that could be reasonably expected.

Applying the Ventsim model to this mine design, the model was run with a total volume of 335 m³/s of which 220 m³/s is exhausted from the primary exhaust fans and 115 m³/s from the Intrepid exhaust fans. The design duty point for each exhaust fan is as follows:

16.3.9.3.1

Primary Exhaust Raise Fans

•

Twin horizontal mount axial fans in a parallel arrangement

•

Design duty point for each fan: 110 m³/s at 3,090 Pa

•

Power consumed during operation: 474 kW per fan

•

Approximate motor size (typical North American supplied motor): 800 hp

16.3.9.3.2

Intrepid Exhaust Fans

•

Twin horizontal mount axial fans in a parallel arrangement

•

Design duty point for each fan: 57.5 m³/s at 2,765 Pa

•

Power consumed during operation: 215 kW per fan

•

Approximate motor size (typical North American supplied motor): 400 hp

An allowance for variable frequency drive (“VFD”) starters for the main fans is included. A VFD allows the electric motors to run at the speed necessary to supply the demand air volume to realize fan energy savings. As a further benefit, any reduction in the air flow volume during winter means that mine air heater energy savings can also be realized.

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16.3.9.4

Mine Air Heating

This study assumes that all intake air entering the mine is at a temperature above freezing point for the following reasons:

•

Protect the health and safety of personnel working or travelling in intake airways.

•

Prevent the freezing of service water and discharge lines.

•

Maintain roadways ice-free and safely trafficable.

•

Prevent rock surface expansion / contraction damage from freezing and thawing of rock joints in the upper parts of the intake airways.

•

Prevent ice build-up in airways that would potentially lead to unsafe conditions, particularly in the primary intake raise hosting the escape-way.

In the colder months the targeted delivery air temperature for the intake air is 2°C. This value was selected on the basis that the air be heated above freezing point while avoiding unnecessary heating costs.

The estimated total heating requirements (combination of both primary intake raise and portal) by month and year are shown in Table 16-25. This table also shows the expected profile of ventilation requirements during the whole of mine life.

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Table 16-25: Estimated Heating Consumption


Airflow	Jan		Feb		Mar		Apr		May		Jun		Jul		Aug		Sep		Oct		Nov		Dec		Total
2017 45 m³/s		114,387		77,830		41,729		7,056		0		0		0		0		0		2,942		36,525		93,451		373,920
2018 135 m³/s		343,160		233,490		125,186		21,169		0		0		0		0		0		8,827		109,576		280,352		1,121,759
2019 225 m³/s		571,933		389,149		208,643		35,281		0		0		0		0		0		14,711		182,627		467,254		1,869,599
2020 260 m³/s		660,900		449,684		241,099		40,769		0		0		0		0		0		17,000		211,036		539,938		2,160,426
2021 305 m³/s		775,287		527,513		282,828		47,825		0		0		0		0		0		19,942		247,561		633,389		2,534,345
Years 2022 to 2026 335 m³/s		851,545		579,400		310,647		52,529		0		0		0		0		0		21,904		271,911		695,689		2,783,625
2027 240 m³/s		610,062		415,093		222,553		37,633		0		0		0		0		0		15,692		194,802		498,404		1,994,239
2028 140 m³/s		355,869		242,137		129,823		21,953		0		0		0		0		0		9,154		113,635		290,736		1,163,306

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16.3.9.5

Auxiliary Ventilation

All work areas in the mine not supplied with a split of fresh air must be ventilated using auxiliary systems. The most effective means for providing airflow to areas without primary airflow is typically with small diameter axial fans combined with low leakage, flexible ducting.

16.3.9.6

Main Decline Development Ventilation

The design criteria for airflow for development of the main decline considered the operation of 1 x 45 Tonne truck (438 kW) and 1 x 7m3 Loader (321 kW). At 0.06 m³/s per diesel kW, 46 m³/s is required for ramp development.

It is planned that this will be accomplished via two auxiliary fan and duct installations in the ramp; one delivering air to the face and the second duct discharging at the remuck. To ensure each fan/duct installation delivers 23 m³/sec to the face, and to account for a small amount of leakage at each duct join (assumed 1.5%), it is estimated that a 110 kW fan is required. This arrangement will deliver the required air for a maximum 600 metres.

Noting that the maximum length of a single heading in the main ramp is expected to be around 1200 m, once a single heading exceeds 600m in length, a second 110 kW fan will need to be placed in series with the original fan for each duct.

16.3.9.7

Level Development and Production Ventilation

During level development and production activities, distances up to 500 m are required to be ventilated using auxiliary systems on each level. The peak individual airflow requirement during development activities will be that for a scoop and a small piece of equipment equal to a total 25 m³/s. At 500 m duct length, a single 110 kW fan will be sufficient, provided good installation and maintenance practices are employed.

16.3.10

Emergency Preparedness

In development of the ventilation strategy for Rainy River, and with due regard to other operational issues, consideration has been given to the potential for mine emergencies. As such, the following criteria have been established:

•

In general, ramps will be in fresh air once developed.

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•

On almost all levels, escape can be either to a ramp or to the escape ladderway in the internal raises.

•

In each ramp, escape may either be up the ramp or down the ramp to a safe area.

•

One permanent 20-person refuge station will be established adjacent to the main decline and access to the 433 ore zone.

•

Three other portable refuge chambers are required for flexibility of location at the most appropriate points in the mine.

•

Whilst the primary means of communication will be by radio, a stench system will be in place for introduction of ethyl mercaptan into both portal and primary fresh air raise concurrently in the event of fire.

•

Fire doors will be located in accordance with legislated requirements and to isolate areas of high fire potential to ensure noxious gases are not distributed through the mine workings.

There are a variety of incidents that will trigger the emergency response plan and/or evacuation plan. Such events may be fire, rock fall, injured personnel or major ventilation equipment breakdown.

In the event that the primary egress (main ramp and portal) is unavailable, a secondary means of egress from the mine must be available to allow evacuation of all underground persons when it is safe to do so.

The primary fresh air raise is designated as the secondary means of egress. An Alimak-Hek elevator system will be installed in the raise with landings at the surface and bottom of the raise adjacent to the main decline, and with an intermediate landing at the exhaust air transfer level. The cage is sized to allow room for a full mine rescue team with apparatus.

For the production stoping blocks, a ladderway is installed in each of the raises located next to main ramps. The raises are sized to afford easy passageway. The route of travel for personnel is to use the ladderway to reach the exhaust air transfer drift and walk to the intermediate Alimak landing.

The primary exhaust raise to surface is for ventilation only and not used as a second means of egress. However, as the Intrepid ore zone is remote from the other ore production areas, the exhaust raise for the zone must be the secondary egress. A ladderway is to be installed in this raise to allow escape to surface.

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A static refuge station will be established adjacent to the main decline and access to the 433 ore zone. It is required to provide refuge during an emergency for 20 persons. Section 10.1 of this report details this station.

The remaining personnel working underground, namely the production development and service crews, are provided refuge by means of three, 12-person mobile self-sufficient rescue chambers. These will be independent of a compressed air supply, with appropriate provisions for safe refuge. They will be located in areas where secondary egress is not, or has not yet been, established, and will be sited relative to the active working areas to be within the average walking pace duration of a personal self-rescuer device.

The primary purposes of fire doors are to prevent noxious gases from reaching workers should they be trapped underground and to prevent fire from spreading as much as possible. Fire doors will be required to isolate the workshop.

16.4

Underground infrastructure

16.4.1

Refuge Stations

There will be three (3) portable and one (1) permanent refuge stations located throughout the mine. The permanent station as shown in Figure 16-39 (main refuge station) will be located at the intersection of the main decline and the ramp to the 433 zone. This station will accommodate 20 persons and will be equipped with an airlock entrance, a battery backup electrical system, an air conditioning unit, a CO²/CO scrubbing unit and emergency supply of first aid, food, water, and oxygen candles. This refuge station will be located in a bay and will be separated from the drift by a concrete wall. Access to the station is through an air-lock system. Emergency air will be supplied by a dedicated air compressor located on surface at the top of the nearby Fresh Air Raise and supplied via piping down the Raise to the refuge station. This refuge station will also service as the main lunchroom. In addition to the main refuge station there will be three smaller, portable refuge stations. They will be located at the Intrepid, ODM and 17 East zones.

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LOGO

Figure 16-39: General Arrangement Refuge Station Plan

16.4.2

Mine Dewatering

Previous hydraulic conductivity work indicated that hydraulic conductivity of deep bedrock increases by a factor of 5 at 1,630 m³/day of inflow. This translates into approximately 18.7 L/s. For this study, AMC and its subconsultant Nordmin, factored this estimate to 28.6 L/s throughout the mine for sizing of sumps and pumps at each station. Mine dewatering for the New Gold Rainy River underground will be handled by a combination of submersible and horizontal centrifugal pumps located throughout the Intrepid Zone, 17 East Zone, 433 Zone and ODM Zone working levels. The pumps will handle ground in-flow and spent drill water via multiple lifts throughout the mine to minimize pump size and power. Maximum drill water volumes reporting to any single sump are estimated at 7.9 L/s. The level sumps are designed to handle 36.6 L/s. A total of nine lift stations are in the current design. The breakdown is as follows:

•

ODM zone: 4 pump stations

•

17 East zone: 3 pump stations

•

Intrepid Zone: 2 pump stations

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These stations will be designed with consistent vertical intervals to maintain similar pump head characteristics, reducing the variety of different pump sizes required. A typical sump will allow for solids settling with an overflow weir to a clear water side before pumping to a holding tank. The holding tank is sized to take several charges from the sump before pumping the tank contents to the next level sump. The sumps in all areas with the exception of Intrepid will report to a pumping station near the workshops. The Intrepid mining zone sumps and the workshop area water will ultimately report to a main sump located in the Intrepid Zone, from where the water is finally pumped out via the portal to surface. Figure 16-40, Figure 16-41 and Figure 16-42 show the typical sump flow diagram and long section and plan view sump general arrangement, respectively.

LOGO

Figure 16-40: Typical Sump Flow Diagram

LOGO

Figure 16-41: Dewatering Level Sump and Pumping

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LOGO

Figure 16-42: General Arrangement Underground Main Sump Layout

16.4.3

Compressed Air

Compressed air for equipment will be supplied by portable and onboard compressors. No central compressed air facility will be constructed and no compressed air reticulation will be required other than for the permanent refuge station. The underground Maintenance and Service Bay area will have a dedicated compressor permanently installed with air lines from the air receiver routed to convenient locations in the shop area. In addition to the permanent compressors, several smaller, portable compressors will be available. An emergency air compressor will be located at the top of the Fresh Air Raise to supply air down to the permanent refuge station in case of an emergency.

16.4.4

Mine Water Supply

Water supply for underground mining processes and dust control will be supplied via a 4” steel line at the portal. The line will continue through the decline following the development to the underground workings. As pressures rise due to increasing static head with depth, Pressure Reducing Valve (“PRV”) stations will be installed to reduce the water supply pressure below 100 psi.

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A typical PRV station as shown in Figure 16-43 will follow the development down until the supply pressure rises above 400 psi, when the station will then be permanently located and another PRV station inserted to continue following the development down. It is expected to have eight (8) permanent PRV stations servicing the Intrepid, 17 East, and ODM levels when the main development for the underground is complete.

LOGO

Figure 16-43: Typical Pressure Reducing Valve (“PRV”) Station

16.4.5

Fuelling and Lubrication

The underground fuel consumption profile for the LOM is shown in Figure 16-44.

Trucks that make regular trips to surface will fuel up mainly above ground in the service area near the portal away from other pit traffic. There will be one main refuelling area by the underground workshops as shown in Figure 16-45, and satellite fuelling stations (Satstats) in the Intrepid, ODM and 17 East zones. Fuel storage and pumping will be accomplished with self-contained satellite portable fuel cubes. The Satstats are equipped with fire suppression system and fuel spillage containment. They can be re-used and re-located to active areas. The lubricants such as grease, hydraulic and break fluids are kept in the main workshop. The main workshops have been designed with fire doors at each end that are normally open. They will shut during a fire emergency in the shops.

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LOGO

Figure 16-44: Underground Fuel Consumption Profile

LOGO

Figure 16-45: General Arrangement – Underground Fuel Station

16.4.6

Workshop Facility

The underground maintenance workshop area, as shown in Figure 16-46, consists of two large bays, each 10 metres wide to accommodate several trucks. The two service bays are joined by a crosscut allowing for forklift and foot traffic to move from one bay to the other without exiting the shop area. One sidewall along the length of each bay will accommodate tool cribs with the crosscut having common short-term storage of oil and greases. One service bay will be

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equipped with a monorail running the length of the bay. The service area will be equipped with a stationary compressor and air lines to power air tools and provide compressed air as needed. Several welding plugs will be distributed throughout this area along with regular electrical plugs to power shop tools. An office, a storage bay and a wash bay will be also located in the maintenance area. The wash bay will have a collection sump with an oil separator installed.

LOGO

Figure 16-46: Underground Maintenance Workshops Layout

16.4.7

Explosives Magazine

Two bays will be provided for the storage of ANFO, explosive accessories such as boosters and detonators and cartridge explosives. The entrance to the storage bays will be controlled with a lockable chain link fence and doors. The bays will contain wooden shelves for storage of the explosives material.

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The arrangements are shown in Figure 16-47 and Figure 16-48.

LOGO

Figure 16-47: ANFO Storage Plan and Sections

LOGO

Figure 16-48: Cap & Powder Magazine Plan and Sections

16.4.8

Underground (CAF Loading Station)

CAF will be required for all stopes with exposed fill and will represent around 60% of all fill placed underground. The remaining 40% will consist of development waste rock directly tipped into stopes. The CAF system will require a surface slurry plant, aggregate production and stockpiles and a loading station underground.

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The underground loading station will have concrete floors, a sump for spilled slurry and a signalling system to control the delivery of slurry from surface. This will be a simple level sensing indicator in the underground tank which will request the delivery of the next batch of slurry from surface.

The loading station will be provided with fixed lighting and frequent cleaning will be required to remove any spillage from the floors and sump areas. Haulage trucks underground will load aggregate from the chute and cement will be sprayed in batches to manage the cement dosing requirement. Figure 16-49 depicts the underground CRF loading station.

LOGO

Figure 16-49: Underground Cemented Rock Fill Loading Station

16.4.9

Communications and Automation

Radio communications are to be established underground by virtue of VHF leaky feeder. VHF-based hand-held radios, fixed vehicle and base station radios will have access to six separate channels. A head end unit will serve as the master repeater relaying voice channels down the leaky feeder media into the drifts. The head end unit will be readily available to connect surface antennas and equipment as necessary.

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One 48 core fibre optic cable will serve IT communications, VOIP and PLC systems. Fibre optic patch panels and fibre to Ethernet switches will be placed in key locations throughout the mine to convert the fibre network to copper for integration with necessary equipment.

Personnel tracking (should it be required) will be accomplished by virtue of a WiFi RFID tag system. Vehicles will also contain RFID tags. The system will be integrated into Impact Software, a browser based tracking and reporting application, allowing operators and mine controllers to monitor, track and allocate personnel and resources. The RFID system can also be integrated into a ventilation-on-demand (“VOD”) system.

Fixed underground monitoring and control will be by virtue of a PLC system. The monitoring and control of underground systems including but not limited to: ventilation, pumping and power monitoring will be by virtue of remote PLC racks placed near the equipment as necessary. The remote racks will come complete with their own processors and, should the communication link fail, the systems they are controlling will continue to operate.

The remote PLC racks will communicate by virtue of a fibre optic backbone stemming from the portal substation. Fibre cables will branch out through the portal into the underground ramp and out into drifts as required. At substations and at remote PLC locations, fibre to copper switches will be installed bridging the network together.

Blasting will be performed by virtue of a VHF radio controlled Smart Blast system. Smart Blast has the capability to initiate non-electric shock tube as well as standard electric blasting caps. The controller receives confirmation back on all of the remote firing devices. The controller and firing devices may be used anywhere there is leaky feeder coverage. Typically the controller is used in a central blasting location.

The voice over IP (“VOIP”) underground telephone system and IP network communications will also be over the fibre optic backbone stemming from the portal substation. A centralized processor will be installed in a key control location to be determined. Fibre cables will branch out through the portal into the underground ramp and out into the drifts as required. At substations, control rooms, lunch rooms, shops and other key locations fibre to copper power over Ethernet (“POE”) switches will be installed bridging the network together.

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16.4.10

Electrical Distribution

The underground LOM power requirement profile is shown in Figure 16-50.

Electrical power will be supplied to the underground mine from a 27.6 kV overhead line from the main surface substation to the underground mine portal. A 5 MVA 27.6 kV / 13.8 kV unit substation will be installed at the portal, complete with five (5) 13.8 kV secondary breakers. The secondary breakers will feed:

•

13.8 kV to a 500 kVA 13.8 kV / 600 V mine power centre for 600 V loads at the portal and for initial ramp development.

•

13.8 kV underground via parallel redundant feeders. Each feeder will have the capacity to source the required power underground, currently estimated at slightly less than 2.5 MVA.

•

13.8 kV to a 2.5 MVA 13.8 kV / 600 V unit substation at the primary exhaust / fresh air raise area via a short 15 kV overhead line from the portal.

•

13.8 kV to a 500 kVA 13.8 kV / 600 V mine power centre at the intrepid exhaust raise area via a short 15 kV overhead line from the portal. 600 V loads including a 400 HP VFD will be fed from this mine power centre.

A building at the primary exhaust and supply raise will contain an electrical room that will house the 2.5 MVA 13.8 kV / 600 V unit substation complete with secondary breakers; as well as:

•

Two 800 HP VFDs for the primary exhaust fans.

•

Two 150 HP VFDs for the primary supply fans.

•

A 600V MCC c/w automatic transfer switch to supply power to the Alimak, as well as emergency air compressors.

A 250 kW 600V self-contained back-up emergency diesel generator will be installed adjacent to the building complete with integral fuel tank and skin-tight enclosure. The generator will tie in to the automatic transfer switch in the MCC and provide emergency power to the emergency air compressor, the emergency egress elevator and the emergency 600 V MCC in the event of an electrical outage.

The parallel redundant underground feeders will route the length of the main ramp and supply mine power centres for main ramp development and will be alternately connected between feeders allowing for the “leap frogging” of power centres without loss of development power. The

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redundant feeder approach also ensures minimal down time if a cable is damaged, allows for the utilization of additional power if required in the future, prevents blackouts on equipment addition and ensures added safety from the portal substation to the underground. Two (2) 1000 kVA mine power centres have been allotted for main ramp development.

The first permanent underground substation will be established near the Intrepid ramp. A 1000 kVA mine power centre will be installed in the substation along with switchgear with parallel line side connections and sufficient load side breakers to allow for ramp feeder redundancy to continue down the ramp and to the Intrepid development area.

The primary underground substation will be located at the shop / clear sump area. A 2000 kVA 13.8 kV / 600 V unit substation complete with secondary breakers will provide power for the clear sump pumps, shop, and underground development as needed.

Each development area will be fed from the redundant main ramp feeders as mining progresses and each area will be fitted with 13.8 kV and 600V distribution to suit mine development / mining. Portable substations will follow the development to provide 600V power from the 13.8 kV line to the mining equipment, fans, and pumps up to a distance of 300 metres away from the substation.

Figure 16-50 illustrates the projected power requirement profile over the LOM.

LOGO

Figure 16-50: Underground Power Requirement Profile

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16.5

Underground Manpower Requirements

16.5.1

Manpower

During the first three (3) years of underground activity, a contract labour force has been assumed. During months 30 through 36, the labour force will transition from contractor to New Gold personnel resources.

16.5.2

Schedule

Hourly labour will generally work a one-week-on, one-week-off, rotation schedule, whereas technical staff and supervision will largely work a normal five-days-on, two-days-off working week. The working time per day is based on two, 10-hour shifts per day. This will allow up to two hours for blasting fumes to clear between shifts, although in practice, less fume clearance time will be required such that mine access may be regained in less than an hour.

AMC estimates the effective working time per shift during production operations to be 7.25 hours in consideration of travel time, daily safety briefs, and pre-start safety checks.

16.5.3

Organization

The underground mining team will be organized into operational groups consisting of supervision, technical support, maintenance and operations. Table 16-26 shows the projected manpower loading to sustain 1,500 tpd of production once the ramp-up period has been completed.

Table 16-26: Underground Manpower Loading At Full Production


Engineering and Geology	# of People
Chief Mine Engineer		1
Senior Mine Engineer		1
Mine Engineers		1
Surveyors		4
Mine Technicians		1
Mine Geologists		2
Geological Technicians		4
Ground Control Technician		1
Ground Control Engineer		1

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Mine Supervision
Mine Superintendent			1
Mine Captain			2
Mine Clerk			1
Mine Operations
Mine Shift Supervisors			8
Jumbo Operators			8
LH Drill Operators			8
LHD Operators			16
Truck Drivers			24
Trainer (equip & safety)			2
Diamond Drillers			2
Blasters			16
Bolters & Ground Support			8
Grader Operator			2
General Labourers			8
Services			12
Backfill Plant Operators			4
Mine Maintenance
Maintenance & Electrical Superintendent			1
Maintenance Planner			1
UG Warehouse Person			2
Mechanical Foreman			1
Welders			2
Mechanics			16
Electrical Foreman			1
Electrician			8
Apprentices			2
Labourers			4
Total			176

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Figure 16-51 shows the manpower loading through the mine life. Years -2 and -1 are periods of construction of surface infrastructure and early development of the open pit. Underground activity commences in Year 1. Initial loading will primarily be provided by the contractor, with technical support provided by New Gold. During Year 3, mine forces will transition from contractor to New Gold personnel.

LOGO

Figure 16-51: Manpower Loading By Year

16.6

Combined Production Schedule

Figure 16-52 and Figure 16-53 show the annual concentrator feed summary and gold production. A combined production schedule for both open pit and underground operations is summarized in Table 16-27.

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Figure 16-52: Annual Concentrator Feed

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Figure 16-53: Annual Concentrator Feed

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Table 16-27: Open Pit and Underground Mine Production Schedule


	Ore Milled (OP & UG Combined)												Gold		Silver		OP Direct Proce- ssing (Mine to Mill)						UG Direct Proce- ssing						OP Stockpiled Total						OP Stockpile Reclaim Total						OP Waste Rock		Over- burden		OP S/ R (exc. Stock- pile)		OP S/ R (inc. Stock- pile)
Period	Ton- nes (Mt)		Au (g/t)		Ag (g/t)		Au Recov- ery (%)			Ag Recov- ery (%)			‘000 oz		‘000 oz		Ton- nes (Mt)		Au (g/t)		Ag (g/t)		Ton- nes (Mt)		Au (g/t)		Ag (g/t)		Ton- nes (Mt)		Au (g/t)		Ag (g/t)		Ton- nes (Mt)		Au (g/t)		Ag (g/t)		Ton- nes (Mt)		Ton- nes (Mt)		OP S/ R (exc. Stock- pile)		OP S/ R (inc. Stock- pile)
2015(Y-2)																														0.16		0.71		1.84								8.8		13.4
2016(Y-1) ¹		0.38		1.50		3.00								17		24		0.38		1.50		3.00								0.75		0.45		1.76								12.7		11.9
2017(Y1)		7.54		1.44		2.13		91.9	%		65.5	%		322		337		7.54		1.44		2.13								6.83		0.42		1.56								26.7		12.1		5.15		6.06
2018(Y2)		7.67		1.48		2.00		92.1	%		65.7	%		337		323		7.65		1.48		1.95		0.02		4.65		24.79		4.30		0.41		1.46								42.7		10.4		6.95		7.51
2019(Y3)		7.67		1.45		3.80		92.0	%		63.1	%		328		591		7.47		1.37		3.03		0.20		4.48		32.69		6.31		0.37		1.91								45.4		9.0		7.29		8.13
2020(Y4)		7.66		1.43		3.56		91.9	%		63.5	%		324		556		7.43		1.31		2.65		0.24		5.13		31.94		4.61		0.41		1.93								44.9		11.8		7.63		8.25
2021(Y5)		7.67		1.38		4.92		91.7	%		61.5	%		312		746		7.24		1.14		4.59		0.42		5.46		10.65		5.03		0.38		2.94								47.6		4.9		7.25		7.95
2022(Y6)		7.67		1.46		4.27		92.0	%		62.4	%		332		657		7.11		1.17		4.18		0.55		5.31		5.41		5.91		0.35		2.43								44.3		0.0		6.23		7.06
2023(Y7)		7.66		1.50		2.74		92.1	%		64.6	%		340		436		7.11		1.21		2.59		0.55		5.23		4.63		3.09		0.34		1.78								29.8		0.0		4.20		4.63
2024(Y8)		7.66		1.57		2.15		92.3	%		65.5	%		358		347		7.11		1.30		1.88		0.55		5.12		5.58		0.48		0.30		1.20								12.6		0.0		1.77		1.84
2025(Y9)		7.67		1.21		2.04		91.0	%		65.6	%		271		331		3.58		1.41		1.59		0.55		4.93		8.51								3.54		0.42		1.50		2.5		0.0		0.36		0.71
2026(Y10)		7.66		0.68		1.64		87.3	%		66.2	%		147		268								0.54		4.57		4.21								7.12		0.39		1.45
2027(Y11)		7.66		0.60		2.59		86.2	%		64.8	%		127		413								0.44		4.37		14.12								7.21		0.37		1.88
2028(Y12)		7.64		0.40		2.81		82.3	%		64.5	%		81		445								0.12		4.29		19.94								7.52		0.34		2.54
2029(Y13)		7.66		0.33		2.11		80.2	%		65.5	%		66		341																				7.66		0.33		2.11
2030(Y14)		4.41		0.55		2.33		74.2	%		64.8	%		58		214																				4.41		0.55		2.33
2031(Y15)

Grand Total		104.28		1.13		2.81		90.6	%		64.1	%		3,402		6,004		62.62		1.31		2.79		4.19		4.96		10.31		37.46		0.39		1.99		37.46		0.39		1.99		318.18		73.57		3.91		6.85

^1.	Gold recovered during the ramp up period is considered as a credit within the Owner’s Costs. The gold sold is used as a credit based on New Gold’s estimated ramp up processing capabilities, recoveries and expenditures during this period.

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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 general process flowsheet for the Rainy River Project is shown in Figure 17-1.

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Figure 17-1: Whole Rock Leach Process Schematic Diagram

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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 1,500 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		Mtpa		7.7
		tpd		21,000
		t/h		951
Duration of Operation		years		13.6
Average Feed Grade (Open Pit and Underground, Blended, First 8.5 Years Operation)		g/t Au g/t Ag		1.48 3.13
Average Feed Grade (Open pit, excluding stockpile)		g/t Au g/t Ag g/t Au		1.31 2.79 4.96
Average Feed Grade (Underground)		g/t Ag g/t Au		10.31 0.39
Average Feed Grade (Stockpile)		g/t Ag		1.99
Average Feed Grade (Open pit and Underground, blended, life of mine)		g/t Au g/t Ag g/t Au		1.12 2.80 2.50
Design Feed Grade		g/t Ag		6.00
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		2,800
SAG Power Requirements		kWh/t		13.3

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Criterion	Unit		Value
Pebble Crusher P₈₀		mm		13
Number of Ball Mills				1
Ball Mill P₈₀		µm		75
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		7,500
Number of Tanks				7

17.2

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 building site layout representing the electrical substation, process plant, stockpile and the primary crusher is presented in Figure 17-2.

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

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17.2.1

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 140 by 120 m to allow for mine truck circulation and an emergency run of mine rock stockpile area on the south end of the pad .

Open pit mine trucks (220-tonne class) 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 596 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 25-tonne gyratory crusher service hatch 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 layout of the building foundation is based on using mechanically stabilized earth (“MSE”) to minimize capital costs. A mechanically stabilized earth wall will be built so that the conveyor to the coarse ore storage remains aboveground. 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 440-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.

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.

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A maintenance building will be built beside the crusher building for maintenance of major crusher components.

Sump pumps at the reclaim level will be sized to handle all the incoming flows including ground water, service water from operation and the runoff water that will come from the conveyor’s trench.

17.2.2

Crushed Rock Handling and Storage

The gyratory crusher discharge 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 14,240-tonne live capacity and an overall capacity of 71,830 tonnes. Under the stockpile, a reclaim tunnel is installed to recover the stored material. A prefabricated electrical and mechanical room is located adjacent and above the reclaim tunnel.

The crushed rock will be reclaimed by three (3) 2,134 mm wide apron feeders all in operation with the capability to operate two (2) at full production capacity while conducting maintenance. The apron feeders 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.

17.2.3

Processing Plant and Tailings Handling

The main processing building houses the grinding circuit (SAG mill, ball mill, and hydrocyclone), pebble crushing, gravity recovery, CIP, carbon stripping, electrowinning, refining and reagent preparation areas, as well as the tailings pumps, compressors and metallurgical laboratory. Three (3) electrical rooms supply power to the plant. 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.

The general plant layout is shown in Figure 17-3.

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

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17.2.4

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 mm. The SAG mill will be equipped with two (2) 7,500 kW motors (dual pinion) 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 pebbles recycled from the pebble crusher to feed the SAG mill.

The scalping screen undersize, with a T₈₀ of 2,800 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 will 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.

17.2.5

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 that 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 tailings will be returned to the ball mill pump box.

17.2.6

Cyanide Leaching Circuit

A 45 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 61% 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 fed to the leach tanks by slurry pumps and diluted to 50% solids with cyanide bearing process water in the leach circuit feed tank. The thickener overflow will flow into the 17 m diameter process water tank (non cyanide bearing) 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 snow, ice and other weather hazards from damaging the equipment and to facilitate operation and maintenance during winter.

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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 22.7 m to 19.2 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.

17.2.7

Carbon-in-Pulp Circuit, Carbon Stripping and Reactivation

A carousel style CIP circuit was selected. Seven (7) tanks with a volume of 360 m³ will be required, with each tank containing 20 tonnes of carbon. The retention time for each tank will be approximately 17 minutes, with approximately 117 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 7,500 g/t. In normal operation, one (1) tank from the CIP circuit will be emptied and transferred to the stripping circuit every two (2) days however the design of the circuit allows for a transfer every day. 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 elute 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 silver sludge, which will be dried and smelted into doré bars as a final product. The now barren solution discharged from the electrowinning cells will then be returned to the barren solution tank.

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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.

17.2.8

Tailings Management and Cyanide Destruction

The CIP tailings will be pumped to a 45 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 objective of the pre-detox thickener is to recycle residual cyanide from the CIP tailings. The overflow from the thickener will be stored in the cyanide bearing process water tank and used as cyanide bearing process water, predominately for 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 (“SO”) will be added along with oxygen (in the 2 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 sulphuric acid (“H₂SO₄”) produced in the cyanide destruction reaction. The cyanide destruction circuit will require one (1) 14 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 tailings pond. Water from the tailings pond will be reclaimed and pumped into the process as non-cyanide bearing water.

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A single pipeline will be used for tailings pumping. The pipeline will be approximately 6.8 km in length at the beginning of operations, will be expanded to a total length of 10.4 km and will be made of HDPE. There will be a set of standby tailings pumps that can be used when the other pumps are in maintenance or shut down. 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 pipelines 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.

17.2.9

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 that 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.

17.2.10

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.

17.2.11

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.

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17.2.12

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.

17.2.13

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.

17.3

Energy, Water and Consumable Requirements

17.3.1

Energy

The power demand of the process plant will be approximately 44.2 MW, or a total energy consumption of 46.5 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
Primary Crushing (Gyratory)		1.2
Grinding (SAG and Ball Mill)		25.7
Processing (Peripherals and Back end)		12.5
Network Loss (2%)		0.7
Power Demand Subtotal		40.2
Security Factor (10%)		4.0
Process Plant Power Demand²		44.2

^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 Chapter 18 for the whole site power demand.

17.3.2

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 Tailings Management Area (“TMA”) and Mine Rock Pond (“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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Figure 17-4: Process Plant and Tailings Pond 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 870 m³/h.

Fresh water will be obtained from the Water Management Pond (“WMP”) 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.

17.3.3

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 150 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.

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.

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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 (2019). 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.

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 44 m³ reservoir.

The total autonomy of the system will be approximately 140 days.

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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 20 t and stored in a 64 m³pressurized 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. In addition, it will also allow delivery of the liquid sulphur dioxide 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.

The total autonomy of the system will be 10 days.

Copper Sulphate

Copper sulphate (“CuSO₄·5H₂ O”) 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.

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The total autonomy of the system will be approximately seven (7) 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 that 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.

Reagent Consumption

A breakdown of the estimated reagent consumptions per tonne milled and total annual consumption is presented in Table 17-3.

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Table 17-3: Process Plant Reagent Consumption


			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	CuSO₄Ÿ5H₂O	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 media consumptions were estimated using two (2) methods (Bond and Molycop Tools) and validated through discussions with Vendors. Consumptions were estimated on a g/kWh basis and tonnage consumptions were estimated based on the annual power consumptions for the SAG 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

Type

Size
(mm)

Estimated Consumption

Equipment

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

A general layout drawing of the infrastructure, including the process plant, offices, administration building, main electrical substation, truck shop, truck wash facilities, coarse ore storage and crusher building is provided in Appendix F.

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 from aerial photographs showing tree and ground cover. Topsoil removal and grubbing are considered to be carried out 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 perimeter of the overburden and waste rock stockpiles.

Site drainage to support construction works will be achieved with the excavation of drainage ditches bordering building platforms and roads, feeding to 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 overstressing the clay. Excavations deeper than 2 m must 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 and underground piping that are not sitting on rock.

18.1.1

Primary Site and Access Roads

The plant site is very well situated and makes use of existing roads. As such, these onsite roads will provide access to the Tailings Management Area (“TMA”) and the explosives plant. 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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Plant site roads will connect the process plant area to the coarse ore storage and the crusher area. The roads represent a total length of approximately 1,050 m.

TBT was commissioned by Rainy River Resources Ltd. 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.

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 from the new proposed East Access Road.

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 6,000 m. These roads will be built at the start of the Project and will remain in use for the duration of the mine life. Mine haul roads will also be built between the open pit and the tailings dam to haul mine material to the dam.

18.1.3

Geotechnical

AMEC carried out a series of geotechnical drilling campaigns at the open pit, tailings and water management 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 the 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 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. A further five (5) boreholes and six (6) test pits were put down in 2013 to obtain information on a clay borrow material area from within the footprint of the Water Management Pond.

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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) and AMEC (2013k) 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.

18.2

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 220 tonne class 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	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

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Table 18-2: Mining Vehicle Repair Bay Specifications


Specification	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. 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.

18.3

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 in the peak year (Year 5, 2021).

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Table 18-3: Staff Requirements


Total Requirements	Peak Personnel # Total		Office Personnel Requirements
Open Pit Mining		318		33
Underground Mining		168		10
Process Plant		91		11
General and Administrative		29		26

Total:		606		80

18.3.1

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 kitchen area and three (3) meeting rooms.

18.3.2

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) 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.

18.3.3

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.

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18.4

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.

18.5

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.

18.6

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 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.

Explosive Plant and Storage

A pad 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.

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18.7

Electrical and Communication

Tie-Point Switching Station, Power Line, Main Substation, and Site Electrical Distribution. The total power demand of the Project was determined to be approximately 58.4 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).

Table 18-4: Estimated Total Project Power Demand


Area	Power Demand (MW)
Processing Plant		39.4
Site Infrastructure (including reclaim water barge)		6.0
Open Pit Mine		2.1
Underground Mine		4.6
Network Loss (2%)		1.0
Power Demand Subtotal		53.1
Security Factor (10%)		5.3

Total Power Demand		58.4

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 currently connecting Fort Frances and Kenora. For interconnection purposes, a 230 kV tie-point switching station is required. An optional feed at 115 kV was studied during the Feasibility Study but was not chosen because of lower reliability, less flexibility and higher losses.

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 15,000 kW (7,500 kW 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.

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 be driven at 4.16 kV or 600 V feed.

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The open pit mine will require an electrification system at 7.2 kV to feed the hydraulic shovel. A dedicated 27.6 kV overhead line will feed a step-down mine substation located near the open pit blast limit. This mine substation will provide 7.5 MVA of power at 7.2 kV for the electrification of the open pit. The 7.2 kV will first be routed on an overhead line to bring it closer to the mining equipment where it will then change over to an insulated cabling system. The switch from overhead line to insulated cables will be optimally placed to minimize the 7.2 kV mine network and shorten the total length of the 7.2 kV trailing cables, which 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 an extension of the 27.6 kV mine OHL. A pole-mounted vacuum recloser (“VCR”) will be installed on the last pole and will allow a transition to insulated cables at 27.6 kV to feed the underground mine electrical system. After a few years, the 27.6 kV OHL can be extended to reach the ODM west vent shaft and provide power at that point.

18.7.1

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 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.

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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.

The proposed emergency generator fleet for the process buildings and the reclaim water (barge) are presented in Table 18-5.

Table 18-5: Proposed Emergency Generators


Location	MW		Voltage
Crusher Building		0.25		600
Concentrator Building		1.50 2.50		600 4,160
Reclaim Water TMA		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 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).

18.7.2

Communication

A combined fibre optic self-healing loop backbone will interconnect all areas. This loop will use the same poles as the 27.6 kV overhead distribution lines and can transmit voice, video and data on the following systems:

•

Telemetry, data acquisition, and control between the process plant and exterior process equipment;

•

Computer network between all departments;

•

Local telephone services;

•

Computer network for maintenance on all electrical equipment data;

•

Fire Detection;

•

Video Surveillance and Access Control Systems; and

•

Electrical Teleprotection Equipment.

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18.8

Tailings Management Area

The means of tailings deposition and storage is a key element for the operability and long-term closure strategy for the RRP. 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 74 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).

18.8.1

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 that 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.

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Table 18-6 summarizes the design parameters for sizing the starter and ultimate dams.

Table 18-6: Tailings Dam Sizing


Parameter	Stage 1 (Starter Dam)	Stage 2	Stage 3	Stage 4	Stage 5 (Ultimate)
Construction Period	Pre-production	By end of Year 2	By end of Year 5	By end of Year 8	By end of Year 11
Dam crest elevation (masl)	366.5	371.5	374.5	377.5	379.5
Maximum height	10.5 m	15.5 m	18.5 m	21.5 m	23.5 m
Struck-level storage capacity with 2 m freeboard to the crest	10.4 M m³	33.2 m³	50.5 m³	69.5 m³	94.6 M m³

18.8.2

Dam Design

The TMA dams have a very high hazard classification due to the potential environmental and economic consequences of a potential failure, and have therefore been designed for the most severe flood and earthquake criteria. The site has low-to-moderate seismic risk with a 0.096 g horizontal peak ground acceleration for a 10,000-year return period earthquake;

The design criteria adopted are:

•

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

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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 4.6H:1V slope and an appropriately sized mine rock toe berm to reduce the overall slope to 5.5H:1V for stability. The slopes are designed to transition between the two 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 mine waste rock and select clay (overburden) and from open pit development, which are available primarily during the pre-production and early years of operations.

The typical cross sections of the tailings dam are comprised of four (4) primary zones, as shown in Figure 18-1 and Figure 18-2. The legend is presented in Table 18-7:

•

Zone 1 is select overburden from open pit development that is placed in lifts with controlled compaction to act as the water retaining element (core) of the dam;

•

Zone 2 is random rockfill (NPAG or PAG) or overburden placed and compacted by the mine fleet to form the upstream shell of the dam;

•

Zone 3 is a NPAG mine rock that forms the downstream shell of the dam and toe berm for stability (South Dam only); and

•

Zone 4 is select or processed sand filter/drain downstream of the core to protect against cracks or construction defects and as a downstream shell foundation filter;

•

Zone 5 is a select sand and gravel or processed rock transition between the Zone 4 filters and Zone 3 rockfill.

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18.8.3

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 that 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 work, such as flattening the overall slope or increasing the size of the toe stabilization berm, could be implemented within an appropriate timeframe.

Overburden zones within the pit pre-strip area that are suitable for dam core 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

(Refer to Table 18-8 for the construction fill material descriptions)

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

(Refer to Table 18-7 for the construction fill material descriptions)

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Table 18-7: Construction Fill Materials (Figures 18-1 and 18-2)


		Construction Fill Materials
1		Core – select clay
2		Shell – random fill
3		Downstream Shell – clean mine rock
4		Filter – Sand
4a		Fine 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

18.8.4

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 into 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 potable water (West Creek Pond). A constructed wetland with four (4) constructed ponds is to be built downstream of the TMA as part of the site effluent treatment system. The locations of the water management infrastructure are shown in the general site drawing (Appendix F).

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Details of the water balance modelling and the overall site water management can be found in AMEC (2013e).

18.8.5

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 Mine Rock Pond (“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 50% of the mill make-up water is provided from the mine rock pond. This rate was selected to ensure that the pond can be kept in balance year over year in mean annual precipitation conditions. Regulation of water recycled to the process plant will ensure there is adequate storage available to contain the EDF with no discharge to the environment.

The West Creek Pond (“WCP”) is established to divert the West Creek flows from a watershed of 808 ha around the open pit. The WCP also provides water to be treated for potable water supply.

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. The TMAP has significant available storage capacity and can store excess water during wet events, or in the event that the effluent cannot be discharged in the desired quantity due to low receiver flows. For preliminary purposes, the TMA/WMP was sized to contain a target volume of 7.0 Mm³ of 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 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. The WMP will also provide a significant portion of the processing plant water demand. Further details on the WMP are provided in AMEC (2013e).

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The Water Discharge Pond (“WDP”) receives decanted water from the water management pond and runoff from the local catchment area (100 ha) and decants managed flows, at a nominal rate of up to 10,000 m³/d from the WMP, 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 maintain effluent treatment efficiencies and to prevent damage to the constructed wetland due to high flows.

The 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 WDP for nutrient and metal removal. The managed volume discharge will help mitigate flow reduction concerns associated with the Pinewood River and site surface water capture.

A flood protection berm/access road is also proposed south of the open pit 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 and 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.

18.8.6

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 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. 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 into the environment.

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Runoff and seepage collected in ditches along much of the south perimeter of the TMA 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 into 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.

Runoff and seepage from the overburden/NPAG stockpile will be collected by east and west perimeter ditches that report to terminal collection ponds (Sediment Ponds #1 and #2) located along the 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 into the environment. If necessary, coagulants and flocculants could be used to aid the settling of suspended solids.

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Table 18-8: Summary of Water Management Ponds


			Maximum		Environmental Design Flood				Dam
Pond	Discharge Pumped (P) / Decant or Spillway (D) to	Water Requirement	Operating Pond (20-year wet year) (Mm³)		EDF Runoff (Mm³)		Pond Volume including the EDF (Mm³)		Crest Elevation (masl)		Average Height (m)		Length (m)
MRP	To process plant (P)	Process water		2.93		0.31		3.24		362.0		5.4		1,650
WCP	To process plant (P)	Potable and other fresh water needs		0.20		N/A		N/A		364.9		4.0		450
TMAP	To WMP (D)	Decanting for discharge		5.57		0.97		6.45		379.5		—		—
WMP	To the environment (Pinewood River below McCallum Creek; (P) To WDP (D)	Process water, freshwater with excess discharged to the environment		6.64		0.13		6.77		373.0		6.7		3,750
WDP	To Constructed Wetland (D)	Excess discharged to the environment		0.08		0.03		0.112		355.2		1.2		360


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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18.8.7

Process Plant Water Supply – Preparations for Start-up

The primary water reservoir to support process plant start-up will be the WMP that 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 in 2016 at the latest.

For the initial start-up, water will be pumped from the local site watersheds and 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 increase 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, and possibly 2017 under extreme drought conditions. 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


	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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18.8.8

Process Plant Water Supply – Operations

At full production capacity the process plant will require an approximate water input of 21,400 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 contact water from the MRP, as well as from the TMAP and the WMP. Under typical, average annual operating conditions, an estimated 10,145 m³/d would be derived from the MRP; 1,690 m³/d would be derived from the WMP; 9,065 m³/d would be derived from the TMAP; and 500 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 7,447 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 recycling back to the process as ‘freshwater’, with the excess discharged to the Pinewood River, either indirectly through the constructed wetland and the lower reach of Loslo Creek, or directly by pipeline to the Pinewood River just downstream of McCallum Creek.

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 (including direct precipitation and runoff), 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 plant site area. Upstream areas of the Clark Creek watershed will be diverted away from the PAG Mine Rock Stockpile area by means of the Clark Creek Diversion.

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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 in preparation for winter, to reduce water inventory losses to ice cover formation, by decreasing the amount of water going to the process plant from the TMAP and increasing the draw from the MRP.

A dedicated pond, the WCP, will be established in line with West Creek for the purpose of providing compensatory fish habitat, and to supply potable water. 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.

18.9

Water Treatment Design Basis and Operation

As part of the feasibility and assessment process, metallurgical aging tests were conducted. These tests revealed that zinc (Zn) and cadmium (Cd) could potentially be leached from the Rainy River Project tailings at a level that could raise concern for meeting regulatory requirements for subsequent effluent quality. Treatability tests conducted at AMEC’s Pointe-Claire laboratory proved that hydroxide precipitation using lime treatment could effectively reduce Cd and Zn concentrations in Rainy River tailings water to desired levels (AMEC 2013h). AMEC was requested to use the results obtained during this testing along with the site water management plan (AMEC 2013i) to design and cost a water treatment plant (“WTP”) for the site at a feasibility level (AMEC 2013j).

Table 18-10 provides a summary of the major WTP design criteria. The design flow was derived from the requirement to treat 7.4 Mm³ in six (6) months, in order to ensure release of properly aged water from the WMP at key times during the year. The minimum flow was set to 50% of this value and the maximum flow, representing the hydraulic capacity of the system, was set to 120% of the design value. The raw water quality assumed was that used for the laboratory

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testing with Zn and Cd concentrations of 50 and 4.2 µg/L, respectively (AMEC 2013h). The design effluent quality targets were established from Provincial and Federal guidelines: less than 20 µg/L for Zn and less than 0.1 µg/L for Cd. Any other heavy metals that may be present would also be removed to low levels by treating these two elements.

Three (3) treatment scenarios were evaluated:

1.	Treating the water in a WTP located between the tailings pond in the tailings management area (TMA) and the water management pond (WMP);

2.	Treating the water in a WTP located between the WMP and the discharge to the environment; and

3.	Treating the water in the WMP.

A comparison of options determined that a modification to Option 3 (Option 3A) was considered most suitable, which involved converting a small portion of the WMP into a settling pond by building an internal dam and then treating the water in the resulting pond.

18.9.1

Water Treatment Design and Operation

Under the preferred treatment scenario (Option 3A), tailings water from the TMA will be raised to a pH > 11.0 with lime and, metal hydroxides will be allowed to settle out of suspension in the settling pond. The continuous settling pond outflow will be acidified to a pH of 8.0 to ensure that biological activity is maintained within the WMP. Sulphuric acid will be used for neutralization purposes, since the use of CO₂ would instead add unnecessary suspended solids into the system. This approach will confine the solids produced due to lime treatment to the settling pond, therefore limiting the area that will require periodic dredging.

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Table 18-10: Summary of WTP Design Criteria


	Units	Minimum		Design		Maximum
Feed Rate
	m³/h		833		1,667		2,000
	m³/d		20,000		40,000		48,000
	L/min		13,889		27,778		33,333
	US gpm		3,669		7,339		8,807
Retention Times (at specified flow)
Reactor #1 (lime)	min		10.0		5.0		4.2
Settling Pond	days		6.0		3.0		2.5
Reactor #2 (acid)	min		10.0		5.0		4.2
Reagent Consumption
Hydrated Lime – Ca(OH)₂	g/L		0.070		0.090		0.160
Quicklime (CaO)	g/L		0.059		0.076		0.136
Sulphuric Acid (93%)	mL/m³		30.0		40.0		50
Sludge Production
Solids Production	g/L		0.060		0.075		0.100

A process flow diagram, process and instrumentation diagrams (3 drawings), and the general arrangements (2 drawings) were prepared (AMEC 2013j).

In order to ensure that there would be no hydraulic short-circuiting of the pond, the WTP location was set in the northwest corner, opposite and distant from the WMP outlet. Selection of this location also helped minimize dam construction costs. In order to have both the reagent addition points and all major equipment in the same general location the settling pond was designed with a u-shaped configuration. This reduces distances for power supply and facilitates operation and maintenance of process equipment.

Due to the existence of significant depths of clay, it was decided to build the plant on structural rock fill. This allows for the plant to be built entirely within the catchment basin of the settling pond itself. The structural fill will begin at the edge of the WMP dam crest and will be graded towards the settling pond.

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The key sludge production parameters are summarized in Table 18-11. The calculation suggests that each year at design flow approximately 27,000 m3 of sludge would need to be dredged. Dredged sludge will be disposed of in the TMA.

Table 18-11: Expected Production of Sludge at Design Flow


Description	Units	Value
Sludge generation rate	g/L		0.075
Sludge production dry weight basis	Mt/d		3
	Mtpa (180 days of operation per year )		540
Sludge percent solids (typical)	w/w		2	%
Sludge production wet wt basis @ 2% solids	m³/d		150
Sludge disposal rate	m³/yr (180 days of operation per year )		27,000

18.9.2

Water Treatment Plant Opportunities

Listed below are a number of opportunities that could lower capital and/or operating costs and may wish to be considered during detailed design:

•

The use of hydrated lime instead of quicklime will reduce the initial capital cost of the lime preparation system at the expense of higher reagent costs. Quicklime was selected at this stage as it is planned for use at the processing plant.

•

Slurried lime could be pumped into a tanker truck and regularly transported to the WTP as needed due to the relatively low lime consumption and since a lime slaker is already planned for the processing plant.

•

The length of the settling pond internal berm could be reduced by a third without affecting settling the overall performance of the pond and constructed in parallel with the WMP dams.

Doing so will save on the extra fill material provisioned in the current design to allow for the flexibility of raising the settling ponds dams, as needed.

•

The building size was established before defining all the process and electrical needs for the plant. It has been determined now that the building size could be reduced without impacting operational ease.

•

The size of the compacted fill could also be reduced. It is necessary to ensure sufficient space for the reagent truck access and building stability but the size has not been optimized.

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•

Reactor #1 could be moved closer to the WTP building. This could save on the lengths of piping and wiring and would also facilitate operation.

18.10

Fish Compensation Works

18.10.1

Background

The RRP is somewhat unique from an environmental perspective in that there are no lakes located within, or adjacent to, the main site. While limited bait fishing does occur with certain project area streams, the area does not support a significant commercial or recreational fishery. Development of the Rainy River Project site will result in unavoidable harm to fish, fish habitat and infilling of waters frequented by fish, which requires the development and implementation of offsets (compensation) pursuant to the Fisheries Act and Metal Mining Effluent Regulations.

The potential impacts to the aquatic environment and fish habitat are as follows (Figure 18-3):

•

Direct loss or alteration of habitat resulting from the infilling and destruction of portions of creeks in the immediate footprint of the mine due to development of the tailings management area, open pit, overburden and mine rock stockpiles, and other infrastructure elements associated with mine development (road crossings, pipeline crossings and outlets);

•

Alteration of habitats due to the realignment or interception of some site watercourses to accommodate project infrastructure or to collect water for processing plant and other usage; and

•

Potential indirect effects to habitat due to flow reductions in the Pinewood River resulting from creek runoff collection at site, groundwater interception by the mine workings (open pit and underground) and/or direct water taking from the Pinewood River (construction and potentially closure / post-closure phases).

Through a collaborative process initiated in mid-2012 with First Nations, Township of Chapple, as well as the Department of Fisheries and Oceans Canada and the Ministry of Natural Resources a fish habitat offset framework has been developed that is being reviewed by the regulatory authorities in parallel with the Federal / Provincial environmental process. Approvals are anticipated to be obtained by either the Fisheries Act Section 35 Authorization requirements, or the Metal Mining Effluent Regulations Schedule 2 amendment process.

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Figure 18-3: Altered/Displaced Waters Frequented by Fish

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18.10.2

Proposed Offset Measures

The predicted impact not associated with mine waste and as such authorized under Section 35(2) of the Fisheries Act represents approximately or 45,543 m² or 10,148 weighted usable area habitat loss as shown in Table 18-13. Weighted usable area is a combined measure of the quality and quantity of habitat such that fish habitat of different locations and types (pond vs. creek) can be compared. The value is unit-less. The proposed habitat compensation balance provides for a net increase in fisheries habitats in a blended approach, including both like for like habitat replacement and watershed-based habitat improvements.

West Creek (including the minor drainage / tributary east of the plant site known as the stockpile pond) will be diverted to the West Creek Pond, which in turn will outlet at its western margin at the West Creek Diversion. The diversion will be constructed to flow in a north-westerly direction before changing direction and flowing in a south-westerly direction where it will converge with the existing Loslo Creek (Cowser Drain) downstream of the proposed constructed wetland. The West Creek pond will be used as a potable water source for the mine during operation, but will maintain a functional wetted habitat throughout mine life and beyond. The West Creek diversion channel downstream of the West Creek pond will be constructed with frequent pooled habitats to provide fish refuge during periods of reduced or intermittent flow which occurs naturally within the system. The pond outlet channel through the emergency spillway will be constructed with a similar low flow channel and slope as the main diversion channel to maintain fish passage between the constructed diversion channel, the pond and upstream sections of the watercourse.

The tributary associated with the stockpile pond east of the plant site will also be diverted through a similarly constructed channel into the West Creek pond.

The Clark Creek system will be intercepted to avoid the east mine rock stockpile and will be diverted through the Clark Creek Diversion Channel to a tributary of the Pinewood River. The developed impoundment structure and the Clark Creek Pond will create sufficient water elevation to redirect flows into the Clark Creek diversion channel. At the terminus of the Clark Creek diversion channel an additional pond (Teeple Road Pond) will be constructed to provide fisheries offsets, and to moderate peak flows from the Clark Creek system into the Pinewood River Tributary. The diversion and impoundment will be created in the same ecotype as the habitat overlaid from Clark Creek.

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The created ponds (West Creek pond, Clark Creek pond, Teeple Road Pond and stockpile pond) will vary in depth. Deeper sections, 1 m or greater, will ensure abundant overwintering conditions in the entire pond habitat, while providing large shallow littoral areas for greater productivity and wetland attributes. Maximum depth will range from 3 to 3.5 m in the West Creek pond, 1.5 to 2.25 m in the Clark Creek pond and Teeple Road Pond and 3.5 to 4 m in the stockpile pond.

Stakeholders have expressed an interest in seeing watershed based and water quality focused offset measures implemented in the local area to mitigate past non-mining impacts in the area rather than just habitat replacement. Restoration measures could include cattle fencing, offline cattle watering sources, and channel and riparian zone restoration. The proposed strategy would make every effort to compliment and work with existing local programs and initiatives, such as the Rainy River First Nations Watershed Program, and Ministry of Natural Resources District Partnership Programs (stewardship council). This means that the compensation program would be set up to support local groups and efforts with a mechanism to track these contributions.

The like for like habitat replacement is fully considered and costed in the Feasibility Update Study. An allowance has also been made for support for watershed-based habitat improvements currently under discussion.

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Table 18-12: Summary of Proposed Habitat Offset Balance for Schedule 2 Amendment Waterbodies


	Total Area Overprinted (m²)						Weighted Usable Area Overprinted							Total Area of Offset (m²)		Weighted Usability Value		Weighted Useable Area Offset
Mine Feature	Loslo Creek (Cowser Drain)		Marr Creek		Total		Loslo Creek (Cowser Drain)		Marr Creek		Total		Offset Feature	Total Area of Offset (m²)		Weighted Usability Value		Weighted Useable Area Offset
Tailing Management Area		143,344		14,949		158,293		32,895		3,434		36,329	West Creek Diversion Channel and Stockpile Pond Diversion Channel		47,241		0.21		9,921
Constructed Wetland / Water Discharge Pond		47,437		0		47,437		10,941		0		10,941	West Creek Pond		150,089		0.23		34,520
West Mine Rock Stockpile		0		5,514		5,514		0		1,230		1,230	Clark Creek Diversion Channel		8,470		0.22		1,863
Overburden Stockpile		0		1,945		1,945		0		428		428	Clark Creek Pond		30,000		0.23		6,900

Total		186,898		22,408		213,189		42,947		5,092		48,928	Total		235,800				53,204

													Net Gain = 22,611 m²				Net Gain = 4,276 WUA

Table 18-13: Summary of Proposed Habitat Offset Balance for Section 35(2) Authorization Waterbodies


	Total Area Overprinted (m² )								Weighted Usable Area (WUA) Overprinted								Offset Feature	Total Area of Offset (m²)	Weighted Usability Value	Weighted Useable Area Offset
Mine Feature	Marr Creek		West Creek		Clark Creek (Teeple Drain)		Total		Marr Creek		West Creek		Clark Creek (Teeple Drain)		Total		Offset Feature	Total Area of Offset (m²)	Weighted Usability Value	Weighted Useable Area Offset
Clark Creek Diversion, East Mine Rock Stockpile		—		—		21,355		21,355		—		—		4,828		4,828	Cattle fencing, offline watering, riparian and channel restoration	To be determined	To be determined	To be determined
Open Pit		—		17,412		—		17,412		—		3,768		—		3,768
Dam Structures		196		—		227		423		41		—		47		88
Plant Site / Ancillary Facilities		—		2,139		—		2,139		—		447		—		447	Clark Creek Pond at Teeple road Like for Like Habitat	Minimum of 45,543 m²	0.23	Minimum of 10,148
Remnant Channels		4,214		—		—		4,214		1,017		—		—		1,017

Total		4,410		19,551		21,582		45,543		1,058		4,215		4,875		10,148

																Net Result	Blended Approach	Minimum 45,543	Varies	Minimum 10,148 plus Offsite

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19.

MARKET STUDIES AND CONTRACTS

19.1

Market Studies

Neither BBA, nor New Gold has conducted a market study in relation to the gold and silver doré that will be produced by the Rainy River Project. Gold and silver are freely traded commodities on the world market for which there is a steady demand from numerous buyers.

19.2

Commodity Price Projections

Commodity pricing is based on base case metal prices and exchange rates consistent with current consensus estimates.

19.3

Contracts

There are no refining agreements or sales contracts currently in place for the Rainy River Project that are relevant to this Technical Report. New Gold expects that terms contained within any sales contract that could be entered into would be typical of and consistent with standard industry practices and be similar to contracts for the supply of gold elsewhere in the world. In the opinion of the QPs, New Gold will be able to market gold produced from the Project.

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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.

Environmental aspects have figured prominently in the development of the preliminary layouts and designs for the Rainy River 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.

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 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 over time throughout the various project stages. Key stakeholders who have demonstrated an interest in the Rainy River Project include the following categories:

•

Municipal Governments: Townships of Alberton, Chapple, Dawson, Morley and Sioux Narrows – Nestor Falls; Towns of Emo, Fort Frances and 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 and Food, Ministry of Economic Development, Trade and Employment, 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 and Ministry of Transportation.

20.2.2

Aboriginal Communications

The Aboriginal groups initially consulted with and engaged in relation to the Rainy River Project were identified using the following criteria:

•

Direction from the Provincial Crown and Federal Crown;

•

Proximity to the Rainy River Project;

•

Past or current interest in similar projects or developments in the region;

•

Demonstrated previous interest in potential biophysical and socio-economic environmental effects of the Rainy River Project; or

•

Aboriginal groups with traditional lands encompassing the Rainy River Project site and its related proposed infrastructure.

Rainy River Resources requested advice from the Ministry of Northern Development and Mines in 2010 and again in 2011 as to which Aboriginal groups should be engaged regarding the Rainy River Project due to the potential impact of exploration and mine development on Aboriginal or Treaty rights. Following advice provided at the time by the Ministry of Northern Development and Mines, Rainy River Resources engaged nine (9) First Nations, along with the Métis Nation of Ontario, that could be affected by the Rainy River Project (Table 20-1).

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Table 20-1: Local Aboriginal Groups Engaged as Instructed by MNDM, December 2011


First Nation or Métis Group	First Nation Number	Reserves Near Project Site	Distance to Rainy River Project Pit Centroid (km)
Anishinaabeg of Naongashiing (Big Island) First Nation		Big Island Mainland 93 Saug-A-Gaw-Sing 1	35 39
	125

		Agency 1 Couchiching 16A	53 40
Couchiching First Nation	126

Lac La Croix First Nation	127	Neguaguon Lake 25D	142
Mishkosiminiziibiing (Big Grassy River) First Nation	124	Big Grassy River 35 G	28
		Agency 1 Rainy Lake 18C	53 47
Mitaanjigamiing First Nation	133

		Agency 1	53
Naicatchewenin First Nation	128	Rainy Lake 17A	29
		Rainy Lake 17B	19
		Agency 1 Rainy Lake 26A Rainy Lake 26B Rainy Lake 26C	53 79 78 91
Nigigoonsiminikaaning First Nation
	129


		Manitou Rapids 11 Long Sault 12	18 21
Rainy River First Nations	130

Seine River First Nation		Seine River 23A Seine River 23B Sturgeon Falls 23	113
	132		90
			119
Sunset Country Métis

In May 2012, the Provincial government identified changes and considerably expanded the list of Aboriginal groups Rainy River Resources would have to consult or notify about mine development (Table 20-2). The majority of these additional groups are located in the Lake of the Woods area. Rainy River Resources elected to have the Provincial Crown coordinate notification in August of 2012.

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Table 20-2: Aboriginal Groups Identified by the Provincial Government

to be Consulted or Notified, May 2012


First Nation or Métis Group	First Nation Number (as applicable)	Reserve Locations (as applicable)	Distance to Rainy River Project Pit Centroid (km)
Aboriginal Groups to Consult:
Anishinaabeg of Naongashiing (Big Island) First Nation	125	Agency 30 Big Island 31D Big Island 31E Big Island 31F Big Island Mainland 93 Lake of the Woods 31B Lake of the Woods 31C Lake of the Woods 31G Lake of the Woods 31H Naongashing 31A Saug-A-Gaw-Sing 1 Shoal Lake 31J	73 55 54 59 35 96 78 92 58 57 39 104












Mishkosiminiziibiing (Big Grassy River) First Nation	124	Agency 30 Assabaska Big Grassy River 35G Lake of the Woods 35J Naongashing 35A Obabikong 35B	73 35 28 49 57 48
Métis – Rainy River Lake of the Woods Regional Consultation Committee Region #1
Naicatchewenin First Nation	128	Agency 1 Rainy Lake 17A Rainy Lake 17B	53
			29
			19
Naotkamegwanning (Whitefish Bay) First Nation	158	Agency 30 Sabaskong Bay 32C Whitefish Bay 32A	73 41 64



		Yellow Girl Bay 32B	76
		Agency 30	73
		Assabaska	35
Ojibways of Onigaming First Nation	131	Sabaskong Bay 35C	40
		Sabaskong Bay 35D	39
		Sabaskong Bay 35F	34
		Sabaskong Bay 35H	42

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First Nation or Métis Group	First Nation Number (as applicable)	Reserve Locations (as applicable)	Distance to Rainy River Project Pit Centroid (km)
		Manitou Rapids 11 Long Sault 12	18 21
Rainy River First Nations	130

Buffalo Point First Nation	265	Agency 30 Buffalo Point 36 Buffalo Point First Nation 1 Buffalo Point First Nation 2 Buffalo Point First Nation 3 Reed River 36A	73 94 103 98 101 100
Aboriginal Groups to Notify:

Anishinabe of Wauzhushk Onigum First Nation (Rat Portage)	153	Agency 30 Kenora 38B	73 101

Couchiching First Nation	126	Agency 1 Couchiching 16A	53 40

Lac La Croix First Nation	127	Neguaguon Lake 25D	142

Mitaanjigamiing (Stanjikoming) First Nation	133	Agency 1 Rainy Lake 18C	53 47

Nigigoonsiminikaaning (Nicickousemenecaning) First Nation	129	Agency 1 Rainy Lake 26A Rainy Lake 26B Rainy Lake 26C	53 79 78 91
		Agency 30	73

Northwest Angle #33 First Nation	151	Northwest Angle 33B	95
		Whitefish Bay 33A	58
		Agency 30	73
		Big Island 37	67
		Lake of the Woods 34	69
		Lake of the Woods 37	85

Northwest Angle #37 First Nation	152	Lake of the Woods 37B	79
		Northwest Angle 34C	104
		Northwest Angle 34C and 37B	102
		Northwest Angle 37C	102
		Shoal Lake 34B1	104
		Shoal Lake 37A	110
		Whitefish Bay 34A	61

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First Nation or Métis Group	First Nation Number (as applicable)	Reserve Locations (as applicable)	Distance to Rainy River Project Pit Centroid (km)
		Seine River 23A	113
Seine River First Nation	132	Seine River 23B	90
		Sturgeon Falls 23	119

Notes:

Rainy River Resources will also continue to consult and involve the Fort Frances Chiefs Secretariat and Pwi-Di-Goo-Zing-Ne-Yaa-Zhing Advisory Services Tribal organizations.

Rainy River Resources elected to have the Provincial Crown coordinate notification in August of 2012.

Rainy River Resources received guidance from the CEA Agency in September 2012 with regard to Aboriginal engagement (Table 20-3). The preliminary depth of consultation provided to Rainy River Resources by the CEA Agency is intended to take into account the strength of the community’s claim to Aboriginal or Treaty Rights and the seriousness of potential adverse impacts.

Table 20-3: Aboriginal Groups Identified by the Federal Government through the

Results of Preliminary Depth of Consultation, September 2012


Preliminary Consultation Depth		First Nation or Métis Group

High:		Naicatchewenin First Nation Rainy River First Nations Anishinaabeg of Naongashiing First Nation (Big Island) Mishkosiminiziibiing (Big Grassy River) First Nation Ojibways of Onigaming First Nation Naotkamegwanning (Whitefish Bay) First Nation







Moderate:		Métis – Rainy River Lake of the Wood Regional Consultation Committee Region #1

Low:		Mitaanjigamiing (Stanjikoming) First Nation Couchiching First Nation Buffalo Point First Nation Northwest Angle #33 First Nation Northwest Angle #37 First Nation Anishinabe of Wauzhushk Onigum First Nation (Rat Portage) Lac La Croix First Nation Seine River First Nation Nigigoonsiminikaaning (Nicickousemenecaning) First Nation

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An important part of the mine permitting and planning process is proactive engagement with area Aboriginal communities. For the Rainy River Project, this engagement included ensuring potentially affected Aboriginal communities are informed and engaged in the development of the project, responding to their interests and concerns, and continuing to build and maintain positive relationships. This has been and is currently being achieved by creating a forum for dialogue and information exchange (verbal and written) and fostering an ongoing relationship between the potentially affected Aboriginal communities and Rainy River Resources.

20.2.3

Comments on the Project

Through the Federal and Provincial environmental process to date, including meeting all regulatory requirements, as well as the voluntary issuance of a draft Terms of Reference and two (2) draft Environmental Assessment (“EA”) Reports (Version 1 for Aboriginal review and Version 2 for all stakeholders and Aboriginal groups) and other measures; Rainy River Resources has made extra efforts to obtain feedback from stakeholders and Aboriginal groups regarding the Rainy River Project. Rainy River Resources has also committed financial resources to the Aboriginal groups for an independent technical review of the draft EA Report (Version 1). The draft Environmental Assessment Report (Version 1) was specifically issued in order to afford Aboriginal groups additional time for review. This has resulted in a very extensive consultation record.

While Rainy River Resources has received positive feedback, including letters of support regarding the Rainy River Project and its potential to bring opportunity to an economically depressed area, some concerns have been expressed by stakeholders and Aboriginal groups regarding the project. Rainy River Resources has responded and attempted to resolve these issues and concerns by a variety of means, including:

•

Alteration to the Rainy River Project where appropriate;

•

Provision for further information or greater clarity on information already provided;

•

Revision to regulatory documentation; and/or

•

Discussions and meetings with the individuals or groups involved.

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Nonetheless, as with all major industrial developments, a number of concerns will require follow-up. The primary concerns expressed are summarized in Table 20-4, along with the proposed approach to reduce and eventually remove the concern.

Table 20-4: Summary of Concerns and Proposed Approach to Resolve


Outstanding Concern or Issue		Proposed Approach to Reduce and Remove Concern
Stakeholders

Additional engineering detail was requested by Government agencies regarding certain project elements, including closure planning.		To be resolved through ongoing meetings and provision of additional information within environmental approval applications, once additional engineering detail is available.




Potential impacts to surface water and groundwater, quality and quantity through the development of the Rainy River Project.		Effluents released from the Rainy River Project are expected to be consistent with Federal and Provincial regulations and policies for environmental protection. Measures have also been taken to limit project flow effects on the receiver, and groundwater drawdown effects on both the receiver and other potential users. Mitigation of the potential effects will continue to be optimized through the construction and operations phase of the Rainy River Project, including ongoing monitoring to confirm impact predictions summarized in the final EA Report. If appropriate, further design changes will be made to ensure compliance with environmental approvals.










Development and operation of the Rainy River Project is expected to impact local aquatic resources and wildlife, largely through displacement of habitat. While compensation will be made in accordance with regulatory requirements, there is a concern that the effects could be greater than anticipated.		Follow up monitoring is proposed to confirm the predicted impact of the Rainy River Project on the local environment. If monitoring should reveal unexpected effects, or effects significantly greater than predicted, then additional mitigation measures would be identified and implemented as appropriate to address the effect. Reclamation once operations cease will return the site to a naturalized setting that will encourage the return of wildlife to the site.

Outstanding Concern or Issue		Proposed Approach to Reduce and Remove Concern
Aboriginal Groups

Traditional Knowledge/Traditional Land Use (TK / TLU): Concern was expressed regarding the role of TK / TLU in the EA and future project planning, the availability of studies lead by each First Nations community and information sharing.		TK/TLU data has been widely collected for the Rainy River Project, including from the closest communities of Big Grassy River First Nation, Rainy River First Nations and Naicatchewenin First Nation. All TK/TLU sessions were community driven, meaning that the method of data collection was community specific. No TK/TLU data has been identified for the Project area specifically. The majority of the data has been broad and overreaching, which Rainy River Resources (Rainy River Resources) will continue to respect as it serves as the basis for First Nations’ unique relationship to the land. TK/TLU collection will continue; information collected will be appropriately considered for construction, operation and closure phases. For example, Rainy River Resources will further investigate the historical travel corridor and appropriately incorporate any new information that may become available.

		Rainy River Resources will share results of the TK/TLU data sessions in a non-public First Nations forum(s).

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Outstanding Concern or Issue		Proposed Approach to Reduce and Remove Concern

Aquatic Resources: Comments were provided regarding the potential for impact to local water quality and fisheries.		Rainy River Resources will commit to a joint water quality monitoring and reporting program with the area’s First Nations as part of the existing monthly water quality monitoring program that is currently carried out by Rainy River Resources. The program will be funded by Rainy River Resources and form an integral part of the overall environmental management program, as it relates to First Nations traditional knowledge and assurances of maintaining water quality and by extension, aquatic biota protection. The program will be developed jointly with the First Nations in lead-up to the initiation of mine construction.

Communication of Information: The First Nations wish to be kept up to date on the Project, including any potential changes.		Rainy River Resources will continue to communicate closely with First Nations regarding the Project.

Environmental Monitoring: Ensure that First Nations have an active role in monitoring plans and programs.		Rainy River Resources has an open invitation for First Nations to participate in all baseline and environmental monitoring programs, including Whip-poor-will, where appropriate, and to share monitoring results. Rainy River Resources will continue to advise on the opportunity of public forums in order to encourage anyone who’s interested in participating.

Cultural Awareness Training: Provide cultural awareness training for those working at the mine.		All Rainy River Resources staff will undergo cultural awareness training. Temporary contractors will undergo an awareness program as part of the regular induction program when working at the mine.

Lake Sturgeon: Consider obtaining new information on Sturgeon.		Additional information related to Lake Sturgeon and the Rainy River First Nations management program will be added to the Final EA Report. Rainy River Resources has committed to a program of close coordination with Rainy River First Nations in support of the pre-existing First Nations Watershed Program and water quality protection. Company funding will be provided as part of the fisheries compensation program to further water quality enhancement programs for the Pinewood and similar agriculturally-impacted waterways.

Baseline Health Information: The Proponent may wish to contact the Seven Generations School and/or Ministry of Natural Resources to obtain additional information.		Rainy River Resources will reach out to the Seven Generations Education Institute and/or the Ministry of Natural Resources to obtain any additional information on baseline health of animals and fish.

Closure Planning: Describe what mechanisms are in place to deliver a successful closure plan over time, including incorporation of TK and community engagement activities.		First Nations will play an active role in the development of the mine Closure Plan, including development of the monitoring and mitigation programs. While the Closure Plan will be completed prior to construction, Rainy River Resources will consult on significant revisions periodically during operations to ensure incorporation of TK and best management practices.

Wildlife Studies: Investigate whether there will be changes to ungulates (moose, deer).		Monitoring programs targeted at ungulates will be coordinated with First Nations.

First Nation Water Supply: Concern expressed regarding the potential for effects to water supply from the Rainy River Project.		Rainy River Resources would be pleased to assemble a map showing the locations of the closest First Nations community water supply intakes upon receipt of the locations/coordinates.

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Outstanding Concern or Issue		Proposed Approach to Reduce and Remove Concern

First Nation Member Health: The First Nations wish to be kept up to date on the Project, including any potential changes. It is suggested that the Proponent and the First Nations work together through a committee to mitigate any potential social problems with workers staying in nearby villages and camps. The largest issue is the potential for more drugs and alcohol to be brought in and consumed in the area.		While the Draft EA has shown no impact to First Nations or non-Aboriginal people’s health, any new information that has a potential to impact health will be provided to First Nations. Rainy River Resources will work with First Nations to ensure employee overall well-being. Programs to highlight the dangers of drug use combined with drug testing will be implemented.

The Métis Nation of Ontario is in the process of completing a TK / TLU and technical review of the Rainy River Project EA Report.		Rainy River Resources anticipates that, as part of the consultation process, an addendum outlining any additional follow-up programs or agreements may need to be submitted in parallel with the final EA Report review.

20.3

Environmental Studies

20.3.1

Overview

Five (5) years of environmental baseline investigations have been completed in support of the Rainy River 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 Project. The intent of this section is to familiarize the reader with the local setting. Further details, including copies of the majority of the baseline reports, are provided in the EA Report.

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 to best 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 for a comparison to be made of pre-development and post development conditions).

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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. Traditional Knowledge (“TK”) and Traditional Land Use (“TLU”) data are also being collected as part of the ongoing Aboriginal consultation program.

The list below contains the environmental baseline study reports prepared for the Rainy River Project:

•

AMEC. 2011. Rainy River Project, 2011 Wildlife Baseline Study.

•

AMEC. 2012H. Rainy River Project, Aquatic Resources 2011 Baseline Investigation.

•

AMEC. 2012I. Rainy River Project, 2011 Wildlife Baseline Study.

•

AMEC. 2012J. Rainy River Project, 2011 Species at Risk Report.

•

AMEC. 2012K. Rainy River Project, 2012 Terrestrial Baseline Study.

•

AMEC. 2012L. Rainy River Project, 2012 Species at Risk Report.

•

AMEC. 2013m. Rainy River Project, Climate, Air Quality and Sound Baseline Study.

•

AMEC. 2013n. Rainy River Project, Socio-economic Baseline Report.

•

AMEC. 2013o. Rainy River Project, 2012 Aquatic Resources Baseline Report.

•

AMEC. 2013p. Rainy River Project, Hydrogeology Baseline Report.

•

AMEC. 2013q. Rainy River Project, Report on Metal Leaching and Acid Rock Drainage Characterization of Mine Rock and Tailings.

•

AMEC 2013r. Rainy River Project. 2013 Aquatic Resources Baseline Report.

•

AMEC. 2013s. Rainy River Project. 2013 Species at Risk Report.

•

AMEC. 2013t. Rainy River Project. 2013 Terrestrial Baseline Study: Bats.

•

AMEC. 2013u. Rainy River Project. 2013 Fish and Fish Habitat Existing Conditions for Highway 600 Realignment.

•

Klohn Crippen Berger. 2011b. Rainy River Project, Baseline Report 2008-2010.

•

Klohn Crippen Berger. 2011c. Geochemical Baseline Report 2008 to 2010.

•

Klohn Crippen Berger. 2011d. Rainy River Project Species at Risk Baseline Report 2008-2010.

•

Ross Archaeological Research Associates. 2011. Stage 1 Archaeological Assessment, Rainy River Advanced Exploration Project, Richardson Township, District of Rainy River. Prepared.

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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.

•

Woodland Heritage Services Limited. 2013. Stage 2 Archaeological and Cultural Heritage Resource Assessment of Rainy River Resources’ Proposed Mining Site, Richardson Township, in the Chapple Township Municipality, Rainy River District, Ontario. MTCS PIF #P208-037-2012 (in progress).

•

Unterman McPhail Associates. 2013. Cultural Heritage Assessment Report: Cultural Landscapes & Built Heritage Resources, Rainy River Project.

Information from other field investigations / work not formally documented as of the issuance of this document may also have been used in the preparation of the EA and Feasibility Update Report.

20.3.2

Climate, Air Quality and Sound

20.3.2.1

Climate

The nearest Environment Canada climate station to the Rainy River 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 695.7 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 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

There are no large urban centres and industrial sources near the Rainy River Project. Background air quality and sound levels are therefore typical of rural, low-density agricultural regions. Air quality in the Rainy River 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.

Background air quality data were developed for the Rainy River Project site from regional monitoring stations that are able to provide multi-year data for a variety of parameters. All parameters assessed were within the maximum desirable level for the National Ambient Air Quality Objectives. Background particulate matter data were supplemented by KCB onsite monitoring.

20.3.2.3

Sound

Ambient noise surveys have been conducted periodically at the Project site since 2009. Measured daytime 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 showing more typical values in the 35 to 40 A-weighted decibel range with the more distant sites (sites >200 m from roads). Night time sound levels were generally lower, as would be expected.

20.3.3

Physiography, Soils and Geology

20.3.3.1

Physiography

Terrain in the Project site area transitions from the upland bedrock controlled lake areas to the northeast, to the lower-lying to gently undulating terrain to the southwest. The Pinewood River system that drains most of the Project site area occupies a broad lacustrine plain. Lands in the 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 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 that left a variably thick veneer of lacustrine clay over much of the landscape.

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The overburden sequence at the 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 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 that 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 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.

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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.

AMEC is conducting environmental geochemical characterizations of selected samples, representative of the mine rock and overburden in the vicinity of the proposed Rainy River Project open 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 the mine rock indicate 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 mine rock 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 to evaluate the long term metal leaching characteristics of these materials. Preliminary results have identified a potential risk of short term neutral metal leaching (cadmium and zinc) from the tailings. These results suggest that a simple lime treatment system may be required for the tailings management area discharge to the water management pond during operations, but not post-closure.

The results of the mine rock and tailings analyzes indicate a future risk for acid rock drainage from a portion of the Rainy River Project mineral waste, if not appropriately managed. The Rainy River Project design has taken this into account in the operation and closure of the east mine rock stockpile and tailings management area.

20.3.4

Hydrology and Hydrogeology

20.3.4.1

Hydrology

Much of the Project area has been cleared over the years for both timber harvesting and through cattle ranching. Most of the natural drainage systems have been altered near the Rainy River Project site through the development of agricultural drains and ongoing beaver activities.

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Regional hydrological data are available from four (4) Water Survey of Canada stations: two (2) on the Pinewood River and two (2) on the much larger Rainy River. More limited flow data are available for Rainy River Project local creek systems. Data from the Pinewood River at Highway 617 are particularly relevant to the local study area because they are on the same river system (the Pinewood River), data are collected year-round, the station is currently active and the watershed is comparatively small, allowing for direct prorated data derivations for other site area watersheds. 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 Pinewood River is a low gradient system with a watershed of 575 km². Local creek catchments draining to the Pinewood River range in size from less than 10 km² to approximately 25 km². The creeks generally originate in rocky uplands, but also frequently originate from or pass through headwater wetland systems. Peak stream flows occur in the spring, with a secondary smaller peak flow in the late fall. Low flows occur in the late winter, and more variably during the late summer and early fall. The average annual runoff for the Rainy River Project site area is approximately 195 mm.

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 that passes through this northeast area, but will not impact directly on any lakes.

Groundwater base flow to area creeks is limited due to the prevailing clay substrates, such that, in certain years, zero or near zero flows are experienced in both local creeks and the Pinewood River during late winter and late summer periods.

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

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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 Finite-Difference Groundwater Flow Model to estimate:

•

Seepage rate into the proposed open pit and the underground workings;

•

Drawdown in Pleistocene lower granular deposit / shallow bedrock, caused by the mine dewatering;

•

Potential reduction in groundwater discharge to the surface water features; and

•

Inflow from tailings management area and east mine rock stockpile, as well as their potential groundwater pathways.

Based on model applications and sensitivity analyses, the predicted groundwater seepage rates into the open pit and underground workings are expected to be in the order of 2,900 m³/d to 3,900 m³/d at full open pit development. The predicted drawdown cone from dewatering of the open pit is predicted to extend approximately 3 to 4 kilometres in all directions 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 impact of a reduction in groundwater discharge to the Pinewood River is expected to be minor and difficult to measure, given that flow in these sections of the river and its tributaries is dominated by surface water runoff contributions and these features can often be dry. A monitoring network of wells and surface water level stations is proposed to confirm the predicted impact.

Based on the high clay content of the Keewatin Till and the associated glacial Lake Agassiz sediments, the local 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

Twenty (20) surface water sampling stations were established for the Rainy River Project during the period of 2007 through 2010, with 14 of these stations still being active and sampled on an approximate monthly basis. Attempts have been made to position stations upstream and downstream of potential future developments within the limitations of the local drainage systems.

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Since June 2010, water quality samples have been collected at approximate monthly intervals to provide a full seasonal spectrum of data, with samples being analyzed for a broad range of general parameters and metals.

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 that 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 2013 from various upstream and downstream stations and analyzed for a suite of parameters. 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 periodically from 2007 to present from monitoring well and drill holes near to the Project site. Samples were analyzed for a complete suite of parameters and compared with Federal and Provincial objectives and guidelines for the protection of aquatic life, as well as the drinking water standards; recognizing that neither of these criteria are not directly applicable. Groundwater baseline water quality reflects the naturally elevated metal content in the local soil and rock. Results from Municipal and private wells showed generally good water quality, with occasional exceedances of drinking water objectives for some parameters.

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20.3.6

Biological Environment - Existing Conditions

20.3.6.1

Aquatic Resources

The Rainy River Project is somewhat unique from an environmental perspective, in that there are no lakes located within, or adjacent to the main Rainy River Project site. While limited bait fishing does occur within certain project area creeks, the area does not support a significant commercial or recreational fishery. In addition, the creeks present within the Rainy River Project site often encounter zero flow during dry periods.

Multi-season studies of fisheries and aquatic resources have been carried out for the Rainy River Project from 2008 through 2013. In the general vicinity of the Rainy River Project area, the Pinewood River shows typical widths of 10 to 15 m, with wider sections associated with beaver impoundments and drowned oxbows. 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 sparse and contribute little to in-stream habitat / cover for fish. Turbidity is high because of erosion of the clay and silt substrates, and agricultural drainage inputs. Beaver dams are frequent and present periodic obstacles to fish passageways in the local study area.

The smaller creeks / Municipal drains that flow into the Pinewood River (Loslo Creek / Cowser Drain, Marr Creek, West Creek, Clark Creek / Teeple Drain, Tait Creek and Blackhawk Creek) 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 fish 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 (also found in Loslo Creek and Clark Creek). Walleye and Yellow Perch occur further downstream in the Pinewood River, but not in the general area of the Rainy River Project site. Lake Sturgeon, classified Provincially as Threatened (Endangered Species Act) in the area, were not present

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within the Rainy River Project site area and environs during the spring 2013 survey. Three (3) Lake Sturgeon were however caught during the spring of 2013 approximately 27 km downstream of the proposed Rainy River Project open pit. As this was the result of a focussed effort at regulatory request, it potentially indicates a lack of suitable habitat and/or a small spawning population specific to the lower Pinewood River.

Small-bodied fish of several species are abundant within the Pinewood River mainstem, as well as in its tributaries. These tributaries likely provide seasonal refuge from predators and contribute to the overall productivity of the Pinewood River system. All fish species present in the system are spring / early summer spawners.

Area benthic communities exhibit a low-to-moderate abundance, with a relatively poor representation of taxa used as indices for characterization and comparison of benthic invertebrate communities, due to the lack of larger substrate particles and dominant clay-silt conditions.

20.3.6.2

Vegetation Communities

The Rainy River Project and environs occur within the Agassiz Clay Plain Ecoregion that 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 that occupy more 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.

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20.3.6.3

Wildlife

Wildlife surveys were carried out for the Rainy River Project and its immediate environs. These surveys were focused principally on birds due to regulatory requirements, 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 and 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. The Mississippi Flyway passes over the regional study area, but results from KCB 2010 migration surveys indicate that the Rainy River Project site area is not considered an important migratory stopover location, as very low numbers of migrating waterfowl, raptors, shorebirds and songbirds were recorded.

Twenty-two mammal species have been identified in the Rainy River Project environs through direct observation, trapping records or sign. Species of cultural significance protected under the Fish and Wildlife Conservation Act, including game animals (Black Bear, Snowshoe Hare, Moose, Elk and White-tailed Deer), furbearers (Red Squirrel, Beaver, Muskrat, weasels, American Mink, American Marten, Fisher, River Otter, Bobcat, Lynx, wolf and Red Fox) and specially protected mammalian guilds (bats, shrews and chipmunks), have been recorded in the local study area. 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. No salamander species were observed. Twelve species of dragonflies/damselflies were observed, or are known to occur, in or adjacent to the Rainy River Project, three (3) of which are provincially rare (Horned Clubtail, Arrowhead Spiketail and Green-faced Clubtail) but do not require special permit or authorization considerations.

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20.3.6.4

Species at Risk and Critical Habitat

The Species at Risk known to occur in the Rainy River Project environs are listed in Table 20-5. 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. The Species at Risk permitting process is currently in progress.

Table 20-5: Species at Risk Known to Occur in the Rainy River Project Environs


Species Common Name	Endangered Species Act	Species at Risk Act
Birds
Barn Swallow	Threatened	—
Bobolink	Threatened	—
Whip-poor-will	Threatened	Threatened
American White Pelican	Threatened	Not at Risk
Bald Eagle	Special Concern	Not at Risk
Canada Warbler	Special Concern	Threatened
Common Nighthawk	Special Concern	Threatened
Golden-winged Warbler	Special Concern	Threatened
Olive-sided Flycatcher	Special Concern	Threatened
Peregrine Falcon (migrant)	Special Concern	Special Concern
Red-headed Woodpecker	Special Concern	Threatened
Short-eared Owl	Special Concern	Special Concern
Mammals
Little Brown Myotis (bat)	Endangered	—
Northern Myotis (bat)	Endangered	—
Reptiles
Snapping Turtle	Special Concern	Special Concern

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, which has a population of 7,952. Excluding the smaller hamlets, Emo is the closest community to the Rainy River 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.

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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’s, at 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 (3) of the ten (10) major employers involved in forestry manufacturing. The remainder of the major employers are in public (health, education, municipal government) and retail services. In 2006, the participation in the labour force for the Rainy River district was 64.2%. This rate is slightly lower than the province’s (67.1%), with a significantly higher unemployment rate (7.9%) compared with the province’s (6.4%). The Rainy River district had a larger workforce share employed in occupations unique to primary industry and trades, transport and equipment operation and related occupations in 2006, compared to Ontario as a whole. This suggests that the workforce is well positioned to service the Rainy River 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 that 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 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 that Rainy River Resources is currently investigating.

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.

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20.4

Traditional Knowledge (“TK”) and Traditional Land Use (“TLU”)

Opportunities for TK / TLU consultations were offered to the following First Nations between July 2012 and February 2013:

•

Anishinaabeg of Naongashiing First Nation (Big Island);

•

Buffalo Point First Nation;

•

Mishkosiminiziibiing (Big Grassy River) First Nation;

•

Naicatchewenin First Nation;

•

Naotkamegwanning (Whitefish Bay) First Nation;

•

Ojibways of Onigaming First Nation; and

•

Rainy River First Nations.

The following First Nations worked closely with Rainy River Resources to collect TK / TLU information:

•

Mishkosiminiziibiing (Big Grassy River) First Nation;

•

Naicatchewenin First Nation; and

•

Rainy River First Nations.

TK / TLU sessions were held with several of the notification Aboriginal groups, including: Couchiching First Nation, Mitaanjigamiing (Stanjikoming) First Nation, and Seine River First Nation.

The TK / TLU studies were led by Ms. Stacey Jack, Rainy River Resources Community Coordinator. Ms. Jack is a licensed Archaeologist, who, as a resident of the District and member of the Couchiching First Nation, has extensive knowledge of regional history. She has worked extensively with area First Nations over the past 20 years, including a leadership role in the development of the Manitou Mounds National Historic Site that is located approximately 35 km south of the Rainy River Project site. In support of the TK / TLU studies, data sharing agreements were signed with these First Nations to ensure that sensitive information is protected and held strictly between the First Nations and Rainy River Resources.

The Métis Nation of Ontario is in the process of completing a TK / TLU study and technical review of the EA report. Rainy River Resources anticipates that as part of the consultation process with the Métis Nation of Ontario an addenda outlining any follow-up programs or agreements may need to be submitted in parallel with the final EA report review.

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Through these consultations, no traditional activities were identified within the Rainy River Project area. Rainy River Resources has committed to undertaking a joint water quality monitoring and reporting program with the area First Nations as part of the existing monthly water quality monitoring program carried out by Rainy River Resources. The program will be funded by Rainy River Resources and form an integral part of the overall environmental management program as it relates to First Nations TK, and assurances of maintaining water quality and by extension aquatic life protection.

Big Grassy River First Nation undertook a second independent review that was provided to the company on October 18, 2013. The review concluded that additional work with the community was required and Rainy River Resources has committed to continuing the close engagement with the community in support of project development.

20.5

Cultural Heritage Resources

The cultural pre-European contact history of the Rainy River District is similar to that in eastern Manitoba and northern Minnesota, and can be divided into the following generalized temporal and cultural sequences: Late Paleo (circa 9000 to 6000 BC), Shield Archaic (circa 6000 to 500 BC) and Middle / Late Woodland (circa 500 BC to AD 1600), and Historic (circa AD 1600 to present).

Prior to the Stage 2 work associated with the Rainy River Project, there were only two (2) previously recorded archaeological registered sites within 15 km of the Rainy River Project. Little information is known about these sites but they appear to be surface collections of unknown age. During 2012 and 2013, Stage 2 archaeological assessment within the Rainy River Project study area, identified eight (8) pre-contact archaeological sites and six (6) historic sites. As required by regulations, these sites have been registered with the Province and are afforded protection under the Ontario Heritage Act until clearance is obtained from the Ministry of Tourism Culture and Sport.

In regard to the already identified sites, preliminary application of the Ministry of Tourism Culture and Sport Stage 3 criteria indicate that four (4) pre-contact archaeological sites and two (2) historic archaeological sites will require Stage 3 excavations. Further Stage 4 work may also be required depending on the Stage 3 results. Consultation with Aboriginal people will continue throughout the archaeological research.

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Unterman McPhail Associates and Jean Simonton, Heritage Consultant, undertook a survey to identify cultural heritage landscapes and built heritage resources within the study area. Twenty-one (21) sites were identified that could be generally described as: rural landscapes (agriculture), township survey, transportation (roadscape), settlement (hamlet), agricultural farm complexes, residences and recreation (trail). The Minister of Ministry of Tourism, Culture and Sport has not designated any of the cultural heritage resources (such as those listed in Table 1 under Part IV of the Ontario Heritage Act) within or adjacent to the study area. In addition, there are no road bridges listed in the Ontario Heritage Bridge Guideline and no Ontario Heritage Trust easement properties or Federally recognized properties within or adjacent to the study area (Unterman McPhail 2013).

20.6

Environmental Sensitivities

There are a number of Species at Risk known or expected to occur within the Rainy River Project footprint or environs. Avoidance of critical Species at Risk habitat is one of the mitigation measures that has already been incorporated into the Project design to a practical level. Based on information currently available, Provincial Species at Risk Permits, pursuant to requirements of the Endangered Species Act, are required for Whip-poor-will and Bobolink. Rainy River Resources with AMEC’s support have been pursuing these approvals in parallel with the EA process.

There are no Federal lands within the Project footprint and Federal Species at Risk Permits will not be required.

There are no Areas of Natural Scientific and Interest, Environmentally Sensitive Areas, Provincially Significant Wetlands, or Federal or Provincial parks within the Project area.

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20.7

Regulatory Context

20.7.1

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 perspective. AMEC understands that the site is fully compliant with existing environmental approvals and has conducted the exploration with due regard to environmental considerations.

20.7.2

Environmental Approvals Required for Proposed Operations

20.7.2.1

Environmental Assessments

Most mining projects in Canada are reviewed under one or more Environmental Assessment (EA) processes. The EA process is a means of project review whereby project design choices, environmental impact 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 Project requires completion of a Federal EA, 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. As the Project had elements that met these requirements, Rainy River submitted a Project Description to the Canadian Environmental Assessment Agency in August 2012 that was subsequently accepted. Based on the Project Description, the Canadian Environmental Assessment Agency confirmed that a Federal Standard EA is required. The Environmental Impact Statement Guidelines that identify the scope of the EA required for the Project were issued on December 18, 2012.

In addition to needing to meet Federal EA 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 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.

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Several aspects of the Project were anticipated to require completion of Provincial EA 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 a portion of gravel-surfaced Highway 600 to avoid potential land use conflicts.

In parallel with the submission of the Project Description to begin the Federal EA process, Rainy River initiated the Provincial EA 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. A subsequent Amended Proposed Terms of Reference was approved by the Ontario Minister of the Environment on May 15, 2013.

Rainy River Resources has been working closely with the Federal and Provincial approvals agencies to harmonize the Federal and Provincial EA processes and, where possible, align public consultation periods to meet the needs of each Act, while minimizing duplication of effort that can lead to unnecessary project delay. This coordination is directed by the Canada-Ontario Agreement on Environmental Assessment Cooperation, and is led by the Canadian Environmental Assessment Agency and the Provincial Ministry of the Environment. Challenges exist as there is no precedent in Ontario and the two (2) regulatory regimes are very different. In December 2012, the Rainy River Project was selected by the Federal/Provincial Regulatory Reform Working Group as one (1) 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.

The Federally-issued Environmental Impact Statement Guidelines and the Provincially-approved Amended Terms of Reference together set out the framework and requirements for the EA Report. The EA Report is intended to provide Federal authorities with information regarding the

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proposed Rainy River Project in order to assist decision making by the Federal Minister of the Environment regarding the applicability of the Canadian Environmental Assessment Act, 2012. It is also intended to provide sufficient information for the Ontario Minister of the Environment to approve the Rainy River Project pursuant to the Ontario Environmental Assessment Act. The Federal and Provincial government authorities have agreed that a single body of knowledge, will be used for the coordinated EA process.

The EA Report was prepared during 2012 and 2013. At various community and leadership meetings, Rainy River Resources was informed that Aboriginal communities did not have the time nor financial and human resource capacities to adequately review the Rainy River Project EA Report. In order to allow adequate time for the Aboriginal technical review, the draft EA Report (Version 1) was released to thirteen Aboriginal groups on May 17, 2013 eight (8) weeks in advance of the general public and government agencies. Rainy River Resources also committed financial resources to the Aboriginal groups for an independent technical review of the draft EA Report (Version1).

A subsequent draft EA Report (Version 2) was provided for government, Aboriginal group and stakeholder review, and hosted at six (6) public venues to facilitate comments starting on July 19, 2013.

Comments received during the draft EA Report reviews were responded to by Rainy River Resources and, as appropriate, were incorporated into a final EA Report issued on December 2, 2013 for a conformity review, as required by the Canadian Environmental Assessment Act, 2012. It is anticipated that final EA Report (Version 2) will be issued for government, Aboriginal group and stakeholder review, and hosted at public venues to facilitate comments, starting in January 2014.

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 EA process with the Provincial Individual EA process being a new approach to enhancing inter-governmental

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alignment in Ontario. While there is always a certain level of risk related to major project EA approval timing, the Rainy River Project is subject to a greater confidence level than there would have been prior to the promulgation of the Canadian Environmental Assessment Act, 2012.

Construction of the Rainy River Project cannot proceed until the Federal and Provincial EAs have been approved, and the appropriate regulatory approvals (as described below) have been attained for the Project component being constructed. Environmental approvals to initiate construction must follow after and are dependent on the EA approvals.

20.7.2.2

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 20-6.

Table 20-6: Anticipated Federal Environmental Approvals


Permit / License	Responsible Agency	Description
Harmful Alteration, Disruption or Destruction of Fish Habitat Fisheries Act	Fisheries and Oceans Canada	Depending on the sensitivity of fish and fish habitat, authorization may potentially be required for the: • Establishment of the mine rock stockpile(s) and tailings management area; • Re-alignment of Highway 600 and mine access, creek crossings; • In water structures such as for fresh water taking; • Watercourse diversions / re-routing; and/or • Flow reductions in watercourses supporting fisheries.

Review of Works in Navigable Waters Navigable Waters Protection Act	Transport Canada	For alteration of navigable waters, such as through establishment of crossing(s) over Pinewood River (if determined to be a Navigable Water); or others.

Schedule 2 Listing Metal Mining Effluent Regulation 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 Metal Mining Effluent Regulation, pursuant to the Fisheries Act.

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Note that explosives magazines, manufacturing facilities and transportation of explosives require a Federal permit, pursuant to Sections 6 and 7 of the Explosives Act. It is anticipated that facilities and equipment owned by the licensed explosives contracted will be permitted through the contractor.

A large number of Provincial approvals are expected to be required to construct, operate and eventually reclaim the Rainy River Project. Key legislation related to the Rainy River Project includes the: Ontario Water Resources Act, Environmental Protection Act, Endangered Species Act, Mining Act, Lakes and Rivers Improvement Act, Public Lands Act and Planning Act. Table 20-7 provides a listing of the Provincial approvals anticipated to be required or likely to be required for the construction and operation of the Rainy River Project. In some instances, multiple approvals may be needed of the same type.

Table 20-7: Anticipated Provincial Environmental 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).

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Approval/Licence	Agency Responsible	Description
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 Net Benefit Permit 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.

Work Permit / 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.

An application for a Leave-to-Construct will also be required pursuant to the Ontario Energy Board Act for approval to construct the proposed transmission line.

20.8

Preliminary Environmental Impact

A comprehensive assessment of the potential environmental impact of the Rainy River Project, along with proposed mitigation measures and the anticipated significance of the effects, has been developed as part of the EA process. The methodology used has been previously accepted for approval of Ontario mining projects and was approved in the Provincial Terms of Reference.

The effects of the Project on a number of valued ecosystem components (VEC; a particular habitat, an environmental feature, a particular assemblage of plants or animals, a particular species of plant or animal, or an indicator of environmental health) and valued socio-economic components (VSEC; components of the socio-economic environment that are significant in terms of people’s values and quality of life) were assessed, for each project phase.

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For each VEC or VSEC identified, the potential effect is analyzed, mitigation or enhancement proposed and an assessment of significance made. Best professional judgement was used in carrying out the effects analysis, incorporating information from available sources, including opinions and perspectives expressed by the various stakeholders and Aboriginal communities through the EA process. Where appropriate, specific analytical methods and tools have been used to support the effects analysis; including laboratory tests, mass balance calculations, statistical packages and various types of models.

Criteria used to evaluate significance included consideration of magnitude / geographic extent, duration, frequency, and ecological / socio-economic context of each effect, as well as whether or not the effect is likely to occur. The direction of the effect (positive or negative) is also considered for socio-economic effects.

The sections that follow provide an overview of the environmental effects analysis and proposed significance, detailed in the final EA Report.

20.9

Effects Analysis

20.9.1

Air Quality and Sound

Air quality emissions were modelled to predict Rainy River Project site area air quality. The potential effect associated with air emissions is an increase in the airborne concentrations of key pollutants in the vicinity of the Rainy River Project site, with the potential to adversely affect air quality. With the appropriate mitigation, the magnitude and geographic extent of any effects on air quality are considered to be low while the duration of the effect on air quality is medium-term, with emissions to the atmosphere throughout the operational life of the Rainy River Project site. The effects are readily reversible, as the air quality effects will cease once the mining and ore processing activities cease (anticipated 16 year mine life) upon closure and reclamation. The overall effect of air emissions, including fugitive dusts, is therefore considered to be minor, as they are limited in geographic extent, limited in magnitude, and reversible.

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Project-related greenhouse gas emissions will result from onsite fuel combustion and other mining and ore processing activities. The estimated maximum emission from the Rainy River Project is very low, and equivalent to approximately 0.02% of Canada’s 2010 emissions.

Sound emissions will vary over the life of the Rainy River Project from lower levels during construction and early operations phases, and increasing gradually to the projected peak in 2020. Beyond 2020, as the open pit continues to deepen and as the stockpiles provide increased shielding, sound levels will begin to decrease and will decline further once open pit operations cease in about 2026. Sound mitigation measures, such as selection of quieter equipment, are inherent to the current design of the Rainy River Project site and are reflected in the sound model predictions. The modelled sound contours for the Rainy River Project site demonstrate compliance with applicable MOE guidelines.

20.9.2

Streamflow, Aquatic Habitats and Species

The Rainy River Project is somewhat unique from an environmental perspective in that there are no lakes located within or adjacent to the main Rainy River Project site. While limited bait fishing does occur within certain project area creeks, the area does not support a significant commercial or recreational fishery. In addition, the creeks present within the Rainy River Project site often encounter zero flow during dry periods.

Project effects are restricted to the minor creeks in the immediate vicinity of the site, including the Loslo Creek / Cowser Municipal Drain, Marr Creek, West Creek, Clark Creek / Teeple Municipal Drain, and the Pinewood River. Development of the Rainy River Project will impact the local creeks and the Pinewood River due to direct habitat displacement (overprinting) and habitat modifications such as channel re-alignment (creeks only); and more indirect pathways such as flow reductions effluent discharge or a combination of the above (creeks and river). The Pinewood River will not be directly altered by any proposed mining works.

Local creeks expected to be directly overprinted by the mine features, in whole or in part, include from east to west:

•

Clark Creek / Teeple Municipal Drain;

•

West Creek;

•

Marr Creek; and

•

Loslo Creek / Cowser Municipal Drain.

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The remaining upstream portions of these creeks, not overprinted directly by mine facilities or infrastructure, will require flow diversion or interception to avoid the upstream flows from interacting with the developed mine areas.

Potential effects on creek flows and water quality will vary from system to system as various portions of the drainages are overprinted by Project components or incorporated into the Rainy River Project integrated water management system. The tailings management area and all stockpiles will incorporate perimeter ditching to intercept runoff and seepage to enable redirection of the drainage to the Rainy River Project water treatment systems and ensure appropriate water quality standards are met prior to release of excess waters.

For the Pinewood River, the annual change in river flow due to the Rainy River Project development will be a function of the capture of watershed that would otherwise flow directly into the river, less the effect of returning of site excess water being returned to the Pinewood River, either through the constructed wetland or by pipeline further downstream (below the McCallum Creek outlet). The net effect will be to reduce Pinewood River flow as measured below the McCallum Creek outlet by approximately 3.45% during an average flow year during early operation, transitioning to a projected overall net gain in flow in the Pinewood River for an average flow year in later mine life of 0.3%.

A combined aquatic habitat displacement or alteration of approximately 26 ha is anticipated. This will result in a harmful alteration, disruption or destruction of fish habitat. Accordingly, a No Net Loss Plan and compensation strategy to offset unavoidable effects to fish habitats is being developed with associated regulatory agencies in consultation with the local communities. A blended offset strategy of watershed restoration with like for like habitat compensation is proposed. Rainy River Resources has committed to supporting water quality and general habitat improvement activities within the Pinewood River watershed along with development of appropriate compensation efforts focused primarily on the establishment of onsite habitats in ponds and diversions, to offset the Rainy River Project habitat losses. Draft versions of these documents have been provided to Fisheries and Oceans Canada.

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20.9.3

Groundwater

Modelling of the proposed open pit anticipates that the zone of influence, defined by 1 metre of drawdown that will eventually develop from the dewatered open pit, is expected to extend approximately 2.5 to 3.5 km from the edge of the open pit in the base case scenario, by the end of mining.

The predicted reduction in the average groundwater flow contribution to the Pinewood River during mine operations is estimated to be three (3) percent of the mean daily flow for the Pinewood River, as measured below the McCallum Creek outlet. Mitigation measures primarily consist of returning captured groundwater as part of the overall Rainy River Project water discharge to the Pinewood River during the period of mine operations to minimize adverse flow effects to the river, especially during low flow conditions; and accelerating open pit inflow following mine closure, to the extent practicable.

20.9.4

Vegetation and Terrestrial Habitat

The primary forest cover types within the natural environment local study area in terms of area extent are:

•

Hardwood forest (47.6% coverage);

•

Coniferous swamp (18.3% coverage);

•

Coniferous forest (9.9% coverage);

•

Agricultural land (7.7%); and

•

Meadow marsh and shallow marsh (4.6%).

The majority of the Rainy River Project footprint overlaps with the hardwood forest community type (mainly aspen-birch), with an anticipated direct displacement of 1,144 ha of hardwood forest community.

Mitigation measures have been incorporated into site planning, with efforts focused on developing a compact site plan with development, where practical, on lands that have been previously disturbed as a result of past anthropogenic disturbance such as logging or agricultural development.

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20.9.5

Terrestrial and Avian Species

Development of Rainy River Project components will affect both terrestrial and avian species through the direct loss of habitats and sound levels in the local vicinity. Development of the Project will also result in an increased risk of mortality due to collisions with vehicles. Many of the species affected will find suitable habitats adjacent to the Rainy River Project given the homogeneous forest cover of the area and abundance of alternative habitats.

20.9.6

Species at Risk

Rainy River Resources has worked collaboratively with the Ministry of Natural Resources on Species at Risk management planning in support of project permitting since 2010. This has included the funding of a collaborative research study along with the Ministry of Natural Resources and Trent University.

No locally significant plant communities have been identified, although two (2) provincially rare plant species were found.

A number of mitigation measures will be employed as part of Rainy River Project development and operation to reduce potential adverse impacts to the species. These measures include:

•

Placement of project components and development of a compact project footprint, in consultation with the Ministry of Natural Resources, to minimize the number of habitat locations that are affected by project components; and

•

Habitat compensation (for Eastern Whip-poor-will) and active protection of known suitable habitats will encourage use by affected individuals.

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20.9.7

Traditional Land Use

While the Rainy River Project will be situated on primarily private lands, Rainy River Resources is continuing to work with the First Nations and Métis to ensure protection of Aboriginal treaty rights. TK / TLU consultations have not identified traditional activities within the Rainy River Project area by the Aboriginal communities that have participated thus far in studies, including Naicatchewenin First Nation, Rainy River First Nations, as well as Big Grassy River First Nation. Some study participants have stated that the Rainy River Project was not an area of intensive use in the distant past, but it is understood that traditional activities may have taken place there.

As a result of the First Nation independent review of the draft EA Report (Version 1), the Rainy River Resources has stated that it will remain open to working with communities.

20.9.8

Socio-economic

The Rainy River Project has the potential, through the generation of very significant employment and business opportunities, to change or influence the population and demographics of the local and regional communities, principally in a positive manner. These potential changes are particularly important considering recent downturns in the forestry and tourism economies that have plagued the region and resulted in a sustained net loss of residents from the District. Improvements to the employment and business economies will also improve the local and regional tax base. Traffic volumes on local roads and highways will increase with proposed Rainy River Project activities during construction and to a lesser extent during operations and decommissioning. There is also the requirement to re-align Highway 600 and to provide alternate access to Marr Road. The expected traffic changes are within the design capacities of the roads affected.

20.9.9

Human Health

Human health can be potentially affected by air and sound emissions and by treated effluents discharged to surface waters, or seepage into groundwater. No such effect is expected as the Rainy River Project will comply with applicable criteria for emissions that are dominantly health-based.

20.9.10

Cultural Heritage Resources

Construction of the Rainy River Project might affect archaeological sites through the disturbance and/or removal of soils during construction and/or operation that can potentially contain the remains of archaeological sites. Proposed Rainy River Project facilities are expected to overprint a number of cultural heritage resource sites pending final site design during detailed engineering. Avoidance has been possible for several other archaeological sites identified but is not expected to be practical for the remainder.

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The Rainy River Project might also affect late 19th and early 20th century built heritage features. Rainy River Resources has committed to undertaking a mitigation program consisting of an illustrated history of the study area.

20.10

Reclamation Approach

Closure of the Rainy River Project site will be governed by the Ontario Mining Act and its associated Regulations and Codes. The Act requires that a Closure Plan be filed for any mining project before the project is undertaken, and that financial assurance be provided before any substantive development takes place to ensure that funds are in place to carry out the Closure Plan.

The objective of closure is to reclaim the mine site area to a naturalized and productive condition when mining ceases. The terms naturalized and productive are interpreted to mean a reclaimed site without infrastructure (unless otherwise negotiated) that, while different from the existing environment, is capable of supporting plant, wildlife and fish communities; and other applicable land uses. A draft Closure Plan is currently in preparation that is anticipated to be issued for government and Aboriginal comment during the first quarter of 2014.

It is expected that the primary phase of active reclamation at the Rainy River Project will take approximately two (2) years after operations cease. Thereafter, the site will be held in care and maintenance until the open pit is fully flooded. Once the pit is flooded, an additional shorter period of active reclamation will occur to remove associated remaining project elements. Environmental monitoring and potentially effluent quality management will occur in accordance with the Closure Plan prepared and filed pursuant to the Mining Act.

20.10.1

Open Pit and Underground Mine

Both the open pit and underground mine will flood naturally once dewatering activities cease. The open pit will be flooded to create a pit lake either passively through natural groundwater entry and precipitation inputs; or by active enhanced flooding of the open pit, using water pumped from an alternate source such as seasonal fresh water inputs. Flooding of the underground and open pit mine to surface is expected to take approximately 72 years using a moderately enhanced, flooding process. Consultation is proposed to determine the preferred flooding approach.

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Other measures to be taken to reclaim the open pit progressively or at closure may, or are likely to include:

•

Remove all infrastructure and equipment within the open pit and underground mine and clean up any petroleum hydrocarbons and/or explosives;

•

Shape and revegetate overburden pit slopes to a stable condition and to facilitate riparian habitats along the pit lake margins;

•

Block the entrance to the open pit and install a boulder or traditional security fence around the pit perimeter during or following active mining operations to ensure safety while the pit is flooding; and

•

Develop a spillway, if needed, to allow the pit lake to eventually overflow into the Pinewood River.

Entrances to the underground mine will be blocked to ensure long term security.

20.10.2

Stockpiles

Progressive rehabilitation of mine rock and overburden stockpiles will be undertaken, where practical, once the maximum height of each stockpile has been reached and/or as each lift is completed, to minimize the amount of reclamation required at closure. All stockpiles will be re-shaped as necessary and stabilized if needed.

The overburden stockpile will be revegetated progressively, with final stabilization and revegetation occurring after overburden has been extracted for site reclamation.

The west mine rock stockpile will contain non-potentially acid generating mine rock. Acid rock drainage / metal leaching is not of concern, so Rainy River Resources proposes to cover the stockpile with a layer of overburden and revegetate.

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A multi-layered cover is proposed for the east mine rock stockpile as it will contain potentially acid generating mine rock. Encapsulation is proposed with a long term goal of controlling acid rock drainage. The side slopes will be covered progressively by a layer of compacted clay till to shed water, topped by a layer of non-potentially acid generating mine rock to consume oxygen, another layer of compacted clay till, followed by a layer of clay till and a growth media to enable revegetation. The flat portion of the stockpile will have a similar cover, but will not include the lowest layer of clay till. Should a temporarily closure or early closure occur, the cover will be completed to ensure acid rock drainage / metal leaching is properly managed.

Rainy River Resources proposed to process all stockpiled ore during operation, therefore reclamation of the low grade or run of mine (high grade) stockpile should not be required. If necessary, the stockpiles will likely be reclaimed in a manner similar to that proposed for the east mine rock stockpile at early or final closure.

Revegetation will occur through seeding, hydroseeding and/or hand planting of tree seedlings as appropriate, to expedite the colonization by indigenous plant species. Investigations will be completed to determine the feasibility of establishing specific wildlife habitats, such as those that might be used by Species at Risk, following closure. The investigations will also determine whether any amendments are required to the native till (overburden) to improve its suitability to provide a base for revegetation.

20.10.3

Tailings Management Area

The principal concerns associated with closure of the tailings management area are long term slope stability, erosion control, drainage, vegetation cover and appearance, as well as prevention of acid rock drainage and metals leaching from the tailings. The tailings management area development plan currently provides for a water and overburden cover at closure to restrict oxygen contact with the tailings surface. Overflow spillway(s) will be developed or deepened to ensure efficient drainage of excess runoff.

20.10.4

Aggregate Sources

If quarries or pits are developed as aggregate sources during the construction and operation phases, these will be reclaimed according to provincial approvals and standards, which may include natural flooding to create pond features.

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20.10.5

Buildings, Machinery, Equipment and Infrastructure

A dedicated onsite demolition landfill is expected to be developed for the disposal of non-hazardous demolition wastes (such as concrete, steel, wallboard and other inert materials) generated by mine closure. It is expected that this demolition landfill will be developed within the east mine rock stockpile.

Salvageable machinery, equipment and other materials will be dismantled and taken off site for sale or re-use if economically feasible, or cleaned of oil and grease where appropriate and deposited within the onsite demolition landfill. Gearboxes or other equipment containing hydrocarbons that cannot be readily cleaned will be removed from equipment and machinery and trucked offsite for disposal at a licensed facility.

All above grade concrete structures will be broken up and demolished to near grade elevation. Concrete structures and below grade facilities (if any) will be infilled if needed. Affected areas will be contoured, covered with overburden as needed and revegetated.

20.10.6

Petroleum Products, Chemicals and Explosives

All petroleum products and chemicals will ultimately be removed from the site. Empty tanks will be sold as scrap, re-used off site, or cleaned to remove any residual fuel / chemicals and deposited within the demolition landfill.

An environmental site investigation will be conducted at the end of operations or early in the closure phase. Soil found to exceed acceptable criteria will be remediated onsite or transported off site to an approved disposal facility.

Any explosives will be depleted towards the end of operations. Any remaining explosives will be either detonated on site or hauled offsite by an authorized transportation company.

20.10.7

Roads, Pipelines and Power Distribution

Site roads may be scarified when no longer needed to support final reclamation, long term site management and environmental monitoring, assuming they are not required to support some other development on the site. Safety berms, if any, along the perimeter of haul roads will be re-shaped to near grade. Culverts will be removed and roads will be breached at the culvert locations on site to allow for natural drainage.

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Pipelines or pipeline sections will either be sealed and left in place; or purged if needed, dismantled and disposed of in the onsite demolition landfill.

Onsite power distribution lines and associated materials that have no salvage value will be dismantled and deposited in the demolition landfill. Other power equipment and materials will be taken off site for sale or re-use.

20.10.8

Site Drainage and Water Structures

The new alignment for the West Creek will naturalize over the life of the mine and will become the permanent creek channel, unless it is determined during closure planning that directing West Creek through part of its original route and on into the open pit is preferred.

The Clark Creek diversion will remain in place to continue to divert drainage away from the east mine rock stockpile.

The pattern of general site drainage will remain in place at closure, with the exception of the removal of culverts at water crossings during site road reclamation activities. Water intake structure(s) at the Pinewood River (or other waterbodies if any) will be reclaimed by removing any structures and mechanical components for disposal in the demolition landfill.

20.10.9

Waste Management

At the end of reclamation activities, the onsite closure landfill will be capped and revegetated in a manner consistent with the remainder of the site and environmental approval requirements.

20.10.10

Offsite Facilities

Highway 600 will remain in its re-aligned form and will continue to provide local access. The re-aligned gravel-surfaced Highway 600 water crossing will remain in place after mine closure.

The East Access Road constructed to support the Rainy River Project development is expected to remain in place.

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It is expected that the 230 kV transmission line constructed to support the Rainy River Project operations will not be required by other local users and will be removed at closure. The option remains to transfer the transmission line to another owner should demand exist at Rainy River Project closure. Assuming reclamation is required, electrical equipment will be removed and recycled / re-used or disposed of. Poles will be removed or cut at grade, and either re-used or disposed of.

Closure costs have been prepared based on this approach and are included in the sustaining capital and financial analysis sections of this study (Chapter 21 and 22, respectively).

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21.

CAPITAL AND OPERATING COSTS

21.1

Capital Costs – Introduction

The capital and operating costs 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. The capital cost estimate related to the open pit mine, concentrator and site infrastructure was developed by BBA and Merit Consultants. AMC Consultants estimated the costs related to the underground mining operation which were included only in the sustaining capital costs. AMEC provided the tailings and water management area material quantities, the water treatment plant costs and site closure costs. The site closure costs and water treatment costs are included only in the sustaining capital costs. BBA and Merit Consultants consolidated the cost information to determine the overall Project capital costs.

The overall capital developed in the Feasibility Study meet the American Association of Cost Engineers (“AACE”) Class 3 requirement of an accuracy range between -10% and +15% of the final Project cost. 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.

21.1.1

Assumptions

The capital cost estimate provides a common basis for classifying all types of facilities and processes and primarily classifies cost estimates based on a measurable degree of engineering completion and prices obtained from vendors. In addition, the capital cost estimate was conducted on the following assumptions:

•

Reflects general accepted practices in the cost engineering profession;

•

Assumes that contracts will be awarded to reputable contractors on a lump sum basis (the contractor is required to submit a total and global price instead of bidding on individual items) and an open shop environment for every trade except heavy earthwork and mining pre-development;

•

Heavy earthwork will be awarded to reputable contractors on a unit cost basis;

•

Open pit mine pre-development will be executed by New Gold;

•

All costs are expressed in constant Q4 2013 Canadian dollars (“CAD”);

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•

Exchange rate to the US Dollar (“USD”) is 0.95 to CAD $1.00;

•

Exchange rate to the EURO is 0.695 to CAD $1.00;

•

Taxes are not included;

•

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 New Gold;

•

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 other than at the reclaim tunnel exit (as established from geotechnical data and from foundation design recommendations by AMEC);

•

Costs will be capitalized until commercial production is achieved (30 days at an average of 60% of production capacity);

•

All excavated material will be disposed on-site; and

•

The Project will adhere to the schedule in Chapter 24.

21.1.2

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.

21.2

Capital Cost Summary

The pre-production (initial) capital cost and contingency for the Project is estimated to be $931.4M, including a $73.3M contingency allocation. Total sustaining capital costs are estimated to be $366.3M. Working capital requirements, to cover the period between commission and first metal sales, and project closure bonding requirements are assumed to be covered by New Gold’s current operations and are not included in the cost estimate. Land acquisition, permitting,

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licensing and project financing costs are included in the Owner’s Costs. The Project capital cost (pre-production and sustaining) summary is outlined in Table 21-1. The capital cost breakdown descriptions are outlined in the subsequent sections.

Table 21-1: Project Capital Cost Summary


Area Description	Pre-Production Capital Costs ($M)		Sustaining Capital Costs ($M)
Overhead Power Line		10.2
Highway 600 Realignment		12.4
Open Pit Overburden Pre-Stripping		39.3
Open Pit Waste Removal and Ore Stockpiling		45.1
Open Pit Mining Equipment		85.1		74.1
Mobile Equipment³				5.2
Underground Mine Development Capital¹				110.5
Underground Mine Sustaining Capital²				138.5
Site Development		117.2		4.0
Process Facilities		312.8
Tailings, Water Management and Treatment		50.0		40.5
Fish Habitat Compensation Costs				2.0
Chapple Township Compensation				2.1
Equipment Salvage Value				(60.5	)
Reclamation and Closure Costs				49.9

Direct Costs Subtotal		672.1

Indirect Costs (excluding Owner’s Cost)		106.3
Owner’s Cost		79.7

Indirect Costs Subtotal		186.0

Contingency		73.3

Total Capital		931.4		366.3

^1.	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.

^2.	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.

^3.	Mobile equipment includes equipment for both the process plant and surface operations.

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21.3

Basis of Estimate

The capital pre-production direct costs can also be divided into the following disciplines: mining, civil/earthworks, concrete, structural, architectural, mechanical, piping, electrical automation / communications and labour. Each discipline is calculated using multiple rates, costs and units for each type of concrete, pipe, equipment, labour, etc. The total costs by discipline are shown in Table 21-2.

Table 21-2: Direct Costs by Discipline


Area Description	Pre-Production Capital Costs ($M)
Mining		169.5
Civil/Earthworks		83.3
Concrete		67.6
Structural		42.1
Architectural		18.1
Mechanical		172.7
Piping		50.0
Electrical		52.2
Automation/Telecommunication		16.5

Total Cost		672.1

Mining

Mining costs include overburden pre-stripping, open pit waste removal, ore stockpiling and mining equipment and total $169.5M. All mining equipment cost estimates are based on a mining fleet owned, operated and maintained by New Gold. Further detail is provided in Sections 21.4.4, 21.4.5 and 21.4.6.

Civil/Earthworks

Earthwork quantities were estimated from drawings, topographical data and geotechnical information. Budgetary quotations were obtained from five (5) different potential contractors. After completion of a commercial analysis, a quote was selected as the basis for this estimate. The total civil/earthworks direct costs is $83.3M. Various total quantities are presented in Table 21-3.

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Table 21-3: Civil/Earthwork Quantities


Quantities	Units	Value
Site Preparation (Clearing & Grubbing)	m³		10,898,000
Site Stripping	m³		542,000
Excavation	m³		1,316,000
Backfill	m³		4,111,000
Slope Protections	m³		268,000
Geotextile Liners	m²		259,000

Concrete

Designs from basic engineering were used to develop the concrete and embedded steel quantities. Unit rates, including formwork and rebar, were estimated from similar projects. The total direct cost for concrete works is $67.6M. Various total quantities are presented in Table 21-4.

Table 21-4: Concrete Quantities


Quantities	Units	Value
Concrete	m³		40,700
Embedded Metals	kg		48,122
Anchor Bolts	kg		147,321
Damproofing, Insulation	m²		3,546

Structural

Preliminary design sketches were used to develop the structural steel quantities. Material was priced from the current steel market values and benchmarked against current projects. The total direct costs for structural and steel works is $42.1M. Various total quantities are presented in Table 21-5.

Table 21-5: Structural Quantities


Quantities	Units	Value
Steel	t		5,603
Decking	m²		19,051
Handrails	m		4,090
Structural Erection	hours		167,990

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Architectural

Siding and roofing quantities were estimated from the General Arrangement (“GA”) drawings. Pricing was based on Merit Consultants’ references on recent data from similar projects. The total direct cost for architectural works is $18.1M. Various total quantities are presented in Table 21-6.

Table 21-6: Architectural Quantities


Quantities	Units	Value
Roofing	m²		11,187
Siding	m²		21,496
Masonry Block Walls	m²		3,743
Architectural Erection	hours		62,568

Mechanical

Full specifications were issued and firm price quotations were obtained from vendors for the following major equipment: gyratory crusher, apron feeders, SAG mill, ball mill, scalping screen, pebble crusher. For process and mechanical equipment packages, equipment datasheets and summary specifications were prepared and budget pricing was obtained from vendors. 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.

A platework list was developed with sizing, weights, and surface areas including lining requirements. A heating, ventilation and air conditioning (“HVAC”) equipment list was developed and fire protection and HVAC material take-offs (“MTOs”) were taken from general arrangement drawings. The total mechanical supply and installation direct cost is $172.7M, total mechanical installation hours are 253,612 and the total equipment pre-purchase cost is $125.4M.

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. The total direct cost for piping works is $50M. Various quantities are shown in Table 21-7.

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Table 21-7: Piping Quantities


Quantities	Units	Value
Carbon Steel (rubber lined)	m		1,899
Carbon Steel (standard)	m		24,224
Ductile Iron	m		2,158
Fiberglass Reinforced Plastic	m		19
High Density Polyethylene	m		42,289
Polyvinylchloride	m		2,146
Stainless Steel	m		4,396
Piping Insulation	m		19,222
Total Small Bore Piping	m		21,895
Total Large Bore Piping	m		55,236
Total Piping and Insulation Installed	hours		260,100

Electrical

An equipment list, including capacities and sizing, was developed from the single line diagrams. MTO for electrical bulk quantities were derived from cable schedules and runs, including cable tray routing layouts. Datasheets were prepared for all major electrical equipment and components and budget pricing was obtained from vendors. For less costly electrical equipment, BBA’s historical cost data was used. The total direct cost for electrical works is $52.2M. Various quantities are shown in Table 21-8.

Table 21-8: Electrical Quantities


Quantities	Units	Value
Cable	m		108,380
Cable Trays	m		5,877
Total Electrical Installation	hours		103,539

Automation/Telecommunications

A detailed instrumentation list was developed from the process flow diagrams. An allowance for communications for the process buildings is included in the capital cost estimate. The total direct cost for Automation/Telecommunications is $16.5M and total installation hours are 52,126.

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Construction Labour

Each discipline mentioned above includes an amalgamation of costs and rates for different types of labour. The total labour of 2.1 million hours was estimated and represents 42% of the total direct costs.

All estimated costs for labour are based on ten (10) hours per day, seven (7) days per week, for a total of 70 hours worked per week. Employee rotations of three (3) weeks of work and one (1) week of rest is expected. There is no allowance for evening shifts except for the mine pre-development which will be executed by New Gold’s mining crews.

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 qualified contractors. These quotes were developed by the contractors from MTOs generated during the Feasibility Study. After a commercial analysis was conducted, a quote was selected as the pricing basis for the estimate. The crew rate for every work area and work element, including the cost of operation for heavy machinery, was calculated based on the selected quote.

For the other trades, a general contractor was contacted to update the single combined crew rate based on non-building trades union that was developed in the 2013 Feasibility Study. In order to take into account potential shortages of qualified non-building trade tradesmen, a single building trade average rate was developed and combined with the non-building trade rate in a ratio of 80% : 20% (non-building trade, non-union workers : building trade, unions workers).

The average crew rates are built from four (4) components as follows:

1.	The direct cost is calculated from the rates of direct labour, such as: 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 crew, planners, and secretaries;

3.	Living Out Allowance estimated at $13.75 per man-hour;

4.	Contractor construction equipment costs are estimated at 15% of the total crew rate for mechanical and piping trades, and at 12% for other trades.

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For non-building trade union workers, 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 hour week, with 40 hours at regular rate and 33.5 hours at overtime rate (1.5 times 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 and indirect personnel are estimated to use respectively 73% and 27% of the construction effort, yielding a rate of $129/hour.

For building trade union workers, the direct and indirect cost components are calculated on the same rotation of 73.5 hours per week and 40 hours at regular rate; however, an overtime multiplier of 2 is applied to the remaining 33.5 hours. Again, this includes provisions for small tools, insurance, overhead and profit, travel allowance, and retention premium, yielding a rate of $149/hour.

Both hourly labour rates were combined in the 80% : 20% ratio (non-building trade, non-union workers: building trade, unions workers) to obtain an average rate of $132/hour.

Productivity

Labour productivity is often the greatest risk factor, followed closely by uncertainties 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 at a given time and place.

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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 21-9:

Table 21-9: Labour Productivity Factors


Calculated Productivity Loss Factor
Trade	Factor
Earthwork		1.00
Concrete		1.15
Structural Steel		1.20
Architectural		1.25
Mechanical		1.12
Piping		1.14
Electrical		1.45
Automation/Telecom		1.45

The productivity loss factors are applied on the installation direct labour hours and evaluated from the following first principles:

•

Earthworks is based on multiple contractor quotes including all the appropriate factors;

•

Concrete is based on contractor quotes including all appropriate factors;

•

Structural steel and architectural are based on BBA’s and Merit’s experience for similar projects;

•

Mechanical and piping are based on highly skilled trades personnel and most works to be performed indoors in enclosed areas and buildings; and

•

Electrical and automation are based on National Electrical Contractors Association (“NECA”) Class 4 electrical labour installation manhours with high voltage works outdoors in elevated work situations.

Winter conditions expected between December 1^st and March 31^st were taken into consideration within the aforementioned productivity factors and were also considered in the first year for civil, concrete and steel works, as indicated in the Project Execution Schedule presented in Chapter 24 of this Study.

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21.4

Pre-Production Capital Costs

21.4.1

Introduction

The pre-production costs total $931M, with $672M in direct costs, $186M in indirect costs and $73M in contingency. The pre-production costs represent the expected expenditure incurred during the pre-production period (2015 and 2016).

21.4.2

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 and is included in the capital cost estimate. The plant will be fed through a new 230 kV power line, 18 km long and will be connected to an existing 230 kV line on Hydro One’s provincial grid. A conceptual line routing has been established and included in the Environmental permitting documentation. The right-of-way of this route falls within the claims owned by New Gold. A more precise routing will be achieved during the design phase of the Turnkey contract that will be awarded for the design and building of the power line. The estimated total is $10.2M CAD including the contractor cost, commissioning, connection cost recovery agreement and other consulting fees and expenses. The total is exclusive of property costs and permits that have been included in the Owner’s Costs. The substation, metering, metering communications and commissioning of metering communications or protection and control systems are excluded from the total and part of the substation cost that is included in the site development total.

21.4.3

Highway 600 Re-alignment

TBT Engineering provided engineering input for the Highway 600 realignment to meet the Ministry of Transportation requirements. Approximately 4.9 km of existing local roads require upgrading and a 6.3 km section of the highway needs to be rebuilt as it passes through the Project site. The cost of the Highway 600 realignment was estimated by Merit based on bulk earthwork quantities provided by TBT Engineering including, rock supply, backfill quantities, over 200,000m³of excavated material and 0.3 km² of clearing and grubbing. The estimated total is $12.4M and includes engineering, contract administration, clearing, grubbing, earth and rock excavation.

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21.4.4

Open Pit Overburden Pre-Stripping, Waste Removal and Ore Stockpiling

The open pit mining pre-production costs for waste removal and ore stockpiling are estimated to be $45.1M and the pre-production costs for overburden pre-stripping are estimated to be $39.3M. The initial cost during the pre-production phase totals $84.4M. More specifically, the costs include:

•

Equipment operating costs;

•

Fuel costs;

•

Blasting costs, including the supplier service and maintenance fees;

•

Hourly labour (e.g. mining fleet operators) and salaried staff costs for production, maintenance, engineering and the geology department; and

•

Other costs include dewatering, pre-split, consultant fees, software licence, etc.

21.4.5

Open Pit Mining Equipment

The open pit equipment pre-production capital cost is $85.1M as presented in Table 21-10. The total cost includes equipment pre-purchased in 2014, 2015 and 2016 and a 10% down payment of pre-purchased equipment to be delivered in 2017.

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Table 21-10: Open Pit Mine Equipment Capital Cost


	Pre-Production Period
Equipment	2015	2016	Total
Haul Truck Fleet
Haul Truck	6	1	7
Shovel Fleet
Hydraulic Shovel (Diesel)	2		2
Drill Fleet
Blastholes Drill	1	1	2
DTH Drill (Reverse Circulation Sample, Pre-Split)	2		2
Support Fleet
Wheel Loader	1		1
Motor Grader	1	1	2
Track-Dozer	3	1	4
Track-Dozer (dyke construction and site rehabilitation)	1		1
Auxiliary Fleet
Water Truck	1		1
Water Truck	1		1
Fuel/Lube Truck	2		2
Boom Truck	1		1
Wheel Loader (Used)	1		1
Tow Haul Truck (Used)	1		1
Hydraulic Crane, truck-mounted	1		1
Dewatering Pump	1	1	2
Mobile Pump	2		2
Service Truck	1		1
Tire Changer, truck-mounted	1		1
Mini Bus	1		1
Pick-up Truck Crew Cab	6	6	12
Stemming Loader	1		1
Aggregate Plan	1		1
Lighting Tower	4		4
Other Equipment
Geotech Monitoring Equipment	1		1
Dispatch system	1		1
Total Open Pit Mine Equipment Costs			$85.1 M

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Mining equipment quantities and costs have been developed based on the mine plan presented in Chapter 16. Mining equipment costs are based on vendor quotes requested during this Feasibility Study and from BBA’s recently updated database of vendor pricing. Contrary to the 2013 Feasibility Study’s approach of lease-to-own, this updated Feasibility Study has assumed that New Gold will not finance the equipment purchases.

The capital cost for the open pit mining equipment is calculated on the basis of a 10% down payment on award and the balance is paid on delivery. Payments incurred during pre-production Years -2 and -1 are included in the pre-production capital cost.

For this Study, it is assumed that mine maintenance and service facilities will be built during the pre-production period. These installations will include a truck fueling station, a permanent truck wash station with mud settling basins and a permanent truck shop facility for mine equipment maintenance consisting of a six (6) bay garage and shops.

21.4.6

Site Development and Process Facilities

Site development costs and process facilities costs and presented in Table 21-11.

Table 21-11: Site Development and Process Facilities Direct Costs


Area Description	Pre-Production Capital Costs ($M)
Site Utilities – Water		24.0
Site Utilities – Power		28.5
Site Infrastructure and Facilities		64.7

Site Development Total		117.2

Process Facilities – Crushing		68.3
Process Facilities – Grinding		104.2
Process Facilities – Gravity, Thickening, Leaching, CIP		119.6
Process Facilities – Refining		48.1
Process Facilities – Reagents		8.2
Process Facilities – Other		4.4

Process Facilities Total		312.8

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Site development costs total $117.2M and include expenses for site utilities such as fresh and fire water systems, mine rock pond, stockpile pond and sedimentation ponds, control room, unit substation and distribution, etc. Site development also includes costs for site infrastructure and facilities such as administration buildings, laboratories, shops and other various buildings. Site development costs include construction for 6.7 km of the East Access Road. Infrastructure has been estimated by BBA based on the developed site plan.

Process facilities costs total $312.8M and include equipment costs, reagent storage costs, common mechanical services and electrical unit substation costs.

The design of the crusher area, the crushed ore stockpile area and the concentrator area is based on BBA’s basic engineering work undertaken between June 2013 and October 2013. The site plan and GA drawings developed during this period have been used to estimate the MTOs for all building materials and earthwork requirements.

Equipment costs have been estimated using budgetary proposals obtained from vendors for most process equipment. For a number of the major process equipment, firm price quotations were obtained during the basic engineering and incorporated into the estimate. For the remaining process and mechanical equipment packages, equipment datasheets and summary specifications were prepared and budget pricing was obtained from vendors.

Information obtained during the interim basic engineering work has been integrated into a detailed equipment list with equipment sizes, capacities, motor power, etc.

21.4.7

Tailings, Water Management and Treatment

The total cost for the tailings, water management and water treatment is $50M. The pre-production costs include the Tailings Management Area, Water Management Pond, Discharge Pond, constructed wetland, Teeple Road Dam construction, the tailings slurry pipeline and the tailings water reclaim system (including the water discharge line from the Water Management Pond to Pinewood river). The West Creek Dam, Mine Rock Pond, Stockpile Pond, collection ponds, sedimentation ponds and Clark Creek Diversion are captured in the Site Development Costs in Section 21.4.6.

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21.4.8

Indirect Costs

Indirect Costs were estimated jointly by BBA and New Gold and are divided into several categories as shown in Table 21-12.

Table 21-12: Indirect Costs


Indirect Costs	Millions ($ M)
Engineering Procurement and Construction Management (“EPCM”)		50.8
Pre-Operation Verification (“POV”)		3.0
Temporary Construction Facilities & Services		10.3
Health, Safety, Security		3.0
External Consultants & Professional Services (Third Parties)		10.0
Start-up and Commissioning Services		2.6
Freight & Logistics, Off-site Warehousing & Handling		13.8
Spares Parts		10.1
First Fill		2.9
Owner’s Cost		79.7

Total Indirect Costs		186.2

Engineering Procurement and Construction Management (“EPCM”)

EPCM service costs were developed based on BBA’s reference data for projects of similar size and schedule. Costs were established based on the number of required hours per task on a defined scope of deliverables and charged using BBA’s hourly rate per required specialty. The total EPCM cost is $50.8M.

The overall percentage ratio on Directs (excluding mining costs) for EPCM costs are as follows:

•

Engineering 4% of direct costs;

•

Project Management 1.25% of direct costs;

•

Procurement 0.4% of direct costs; and

•

Construction Management 4.5% of direct costs.

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Pre-Operation Verification (“POV”)

POV (pre-commissioning) costs are included in the estimate for the preparation of the detailed commissioning and process plant start-up procedures.

These costs include the following categories:

•

POV manager salary;

•

Preparation of commissioning plan;

•

Electrical pre-commissioning;

•

Senior minerals engineer salary;

•

Senior mechanical engineer salary;

•

Construction and Pre-commissioning support (programmable logic controller (“PLC”) programmers);

•

Commissioning and Start-up support; and

•

Travel and living expenses.

The total POV cost is $3.0M.

Temporary Construction Facilities & Services

Construction operation costs include the construction and maintenance of temporary worker facilities required during the Project’s construction period. An itemized list with budget allowances was developed by BBA which includes the costs for site distribution of temporary construction power, access roads to the temporary construction facilities, a telecommunication tower and general communications, and other related equipment. A monthly fee of $15,000 over a period of 24 months was included for the construction management team requirements to connect to the Owner’s network (phone, internet, etc.).

Construction workers housing on site was excluded from the capital cost estimate. A provision for Living Out Allowance (“LOA”) and travel costs of $13.75/hour were included instead as part of the labour rate for all trades.

The total cost of temporary construction facilities and services cost is $10.3M.

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External Consultants & Professional Services (Third Parties)

Cost of sub-consultants and other third parties were estimated based on projects of similar size and are included in the Capital Cost Estimate. Third party costs amount to $10M and can be divided into external consultants and professional services.

Third party costs for external consultants were broken down as follows:

•

Outside engineering consultants (undefined): $758,500;

•

Engineering and quality surveillance for the Highway 600 Realignment (based on a cost estimate provided by TBT Engineering): $1.3M;

•

Engineering and quality surveillance for the East Access Road construction: (based on a cost estimate provided by TBT Engineering): $536 K; and

•

Engineering services for water management/geotechnical (based on a cost estimate provided by AMEC): $5.4M.

Third party costs for professional services were broken down as follows:

•

Labour relation management assistance (based on a provision of 200 hours): $37,000;

•

Water treatment testing services (provision): $50,000; and

•

Site surveying services (based on 176 person-weeks of site surveying team): $1.7M.

Start-up and Commissioning Services

Start-up and commissioning costs were estimated based on projects of similar size and include specialist services required to assist New Gold in the final acceptance of the work and plant start-up. Costs also include an allocation of construction manpower resources for miscellaneous commissioning adjustments. The total cost of start-up and commissioning services is $2.6M.

Freight and Logistics, Off-Site Warehousing and Handling

A freight forwarder provided a quote for equipment freight costs. The quote was approximately 6.75% of the equipment purchase cost. It was assumed that most major mechanical and electrical equipment will be stored in an off-site location within approximately a day of travel from the storage area to the site. The total cost of freight and logistics, off-site warehousing and material handling is $13.8M.

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Spare Parts

Costs related to construction and commissioning spare parts were estimated at $10.1M and broken down into the following categories:

•

Capital spares based on a detailed spare parts list: $6.6M (5.4% of total equipment purchase cost);

•

2-year operating spares $2.2M (1.8% of total equipment purchase cost); and

•

Mining fleet spare parts $1.3M (1% of total equipment purchase cost).

First Fill

A detailed list and costs for grinding media, chemicals and lubricants is included in the capital cost estimate. First fills amount to a total cost of $2.9M.

21.4.9

Owner’s Costs

New Gold prepared an itemized list with budget allowances for the Owner’s Costs. Owner’s Costs total $79.7M and 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. Plant operating and maintenance personnel hired during the pre-production period were also included in Owner’s costs. Operating costs are capitalized within the Owner’s costs until commercial production is achieved, which is defined as 30 days at an average of 60% production capacity.

21.4.10

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 Project’s scope, price escalation or currency fluctuations. The value of the contingency, $73.3M, is expected to be spent during the life of the project.

Contingency calculations on discipline work categories and labour totals was performed by the Project’s estimator. The Project’s estimator used knowledge of the actual project engineering completion to date serving as the basis to establish the contingency percentages, following established industry guidelines for an AACE Class 3 estimate. The outcome of the contingency calculation yields an overall percentage of 12% on the total direct and indirect costs. No

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contingency was calculated on the mining and Owner’s costs as these have contingencies built into their totals. Typical contingency for AACE Class 3 estimates range between 10% and 15% of direct and indirect costs, depending on engineering completion.

Material and equipment costs were meticulously developed and provide a high level of confidence in the estimate. Installation productivities were based on past experience and input from contractors. The quantities provided by the engineers are in the expected range for a plant of this size and capacity.

21.5

Sustaining Capital Costs

21.5.1

Introduction

The total sustaining capital cost is $366M. This is the estimated expense required to maintain operations throughout the production life. Sustaining capital costs include open pit mining equipment, surface and process plant mobile equipment, underground mining and development costs, fencing, tailings and water management, compensation costs, reclamation and closure costs.

21.5.2

Open Pit Mine Equipment

Completion of payments in Year 1 (2017) and subsequent equipment purchases are considered to be a sustaining capital cost. A total sustaining cost of $74.1M is expected for open pit mining equipment. The total cost is inclusive of a 10% equipment down payment disbursed one year prior to utilization/commissioning in addition to the remaining payments.

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The sustaining open pit mining equipment are shown in Table 21-13:

Table 21-13: Sustaining Open Pit Mine Equipment


Sustaining Mining Equipment	Production Period
	2017	2018	2019	2020	Total
Haul Truck Fleet
Haul Truck	6	6	2	1	15
Shovel Fleet
Hydraulic Shovel (Electric)	1				1
Drill Fleet
Blastholes Drill	1				1
Support Fleet
Motor Grader	1				1
Track-Dozer	1				1
Wheel-Dozer	1				1
Auxiliary Fleet
Compactor	1				1
Boom Truck	1				1
Wheel Loader (Used)				1	1
Mobile Pump	2				2
Mini Bus	1				1
Stemming Loader				1	1
Cable Reeler (Used)	1				1
Lighting Tower	2				2
Total Open Pit Mine Equipment Costs					$74.1 M

21.5.3

Mobile Equipment

Sustaining capital for the process plant includes an allowance of $5.2M for mobile and shop equipment. Half the allowance will be spent in Year 5 and the balance will be spent in Year 10. The mobile equipment include items for surface work (general and administrative) and process plant such as aerial work platforms, fork lifts, pick-up trucks, service trucks, wheel loader and portable diesel welders.

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21.5.4

Site Development

A provision of $4.0M for project security fencing is included and will be paid in $1.0M installments for the first four (4) years of the mine life (2017-2020).

21.5.5

Underground Mine

The capital costs for the underground mine were estimated based on first principle work-ups of all key mining activities using a detailed cost model developed by AMC. It is estimated that $249.1M in capital will be required to build and sustain the Underground mine. This covers all capital development, mobile and stationary equipment purchases and rebuilds, capitalized pre-production development, and contracted construction. Construction costs include the underground workshop, backfill station, ancillary installations, portal construction, primary electrical distribution, dewatering, and primary ventilation equipment. A contingency representing 8.4% of the total underground capital cost estimate is included.

The underground capital cost estimate is compiled based on the following key points:

•

Budget quotations for all major mobile and fixed plant equipment obtained from vendors;

•

All mine operating costs during the pre-production period (Years 2015 – 2016) are capitalized;

•

Drawings and construction cost estimates for the underground workshop, ancillary installations, portal construction, electrical, dewatering, and ventilation systems have been estimated by Nordmin Engineering;

•

Lateral capital development and capitalized operating expenses are developed from first principle work-ups;

•

Raise development will be executed by a contractor. Unit costs for these activities (including labour) were obtained from contractor budget quotes. Contractor quotes do not include fuel, power, support equipment, and waste removal. These are considered as Owner’s Costs and have been estimated from first principle work-ups;

•

All capital development and mine operating activities during the first three years (2015 –2017) are to be completed by a contract workforce. Labour rates in these years include contractor premiums. All lateral capital development and mine operating activities are to be completed by New Gold in the following years (2018 – 2026); and

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•

The contingency of $20.9M represents 8.4% of the total underground capital cost estimate. This includes a 10% contingency on underground mining costs, 15% contingency on fixed plant purchases and fixed plant sustaining costs, 15% contingency on construction costs, and a 10% contingency on capital indirect costs. Considering the use of actual supplier quotes, no contingency is applied to any purchases of new mobile mining equipment and equipment rebuilds/replacements.

The total underground mining capital requirement is estimated to be $249.1M. Table 21-14 shows the annual capital requirements for the underground mine. The capital requirements are also shown by cost item in Table 21-15 :

Table 21-14: Life of Mine Underground Capital Costs


Year	Annual Capital Costs ($M)
2017		50.7
2018		59.9
2019		30.7
2020		25.7
2021		11.9
2022		9.6
2023		13.1
2024		19.7
2025		16.1
2026		5.6
2027		2.8
2028		3.4

Total		249.1

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Table 21-15 : Underground Capital Cost Breakdown


Item	Total Cost ($M)		Percentage of Total
Capital Development		88.1		35.4	%
Equipment Purchases – Mobile and Fixed Plant		47.5		19.1	%
Sustaining Capital – Plant & Equipment		38.5		15.5	%
Capitalized Pre-Production Development		31.1		12.5	%
Contracted Construction Costs		13.0		5.2	%
Project Indirects		9.7		3.9	%
Contingency		20.9		8.4	%

Total		249.1		100.0	%

Capital development includes the main decline, internal production ramps, ventilation raises, ventilation accesses, main pump station, development for the underground workshop, CAF plant and ancillary installations, i.e., explosive storage facilities, wash bay, and refuge stations. Capital development costs are summarized in Table 21-16.

Table 21-16: Underground Capital Development Costs


Year	Lateral Development ($M)		Vertical Raising ($M)		Total ($M)
2017		3.9		0.0		3.9
2018		13.1		3.3		16.4
2019		12.0		4.1		16.1
2020		14.1		1.3		15.3
2024		6.5		0.1		6.7
2022		4.8		0.0		4.8
2023		5.3		0.1		5.4
2024		7.6		0.2		7.8
2025		7.9		0.3		8.2
2026		3.0		0.0		3.0
2027		0.5		0.0		0.5
2028		0.0		0.0		0.0

Total		78.7		9.5		88.1

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The quantity and total cost of mobile equipment required for the underground mine are summarized in Table 21-17. Equipment purchases are scheduled based on the planned build-up of mine development and production rates. Rebuild and replacement schedules of mobile mine equipment are calculated based on annual equipment operating hours.

Table 21-17: Underground Equipment Purchases – Mobile


Equipment Type	Quantity		Total Cost ($M)
Drill Jumbo, 2-Boom		3		3.70
Drill Longhole		2		2.18
Haulage Truck, 45 Tonne		6		8.95
LHD, 7.2m³with Remote		4		5.84
Bolter		2		1.97
Lubrication Service Truck		2		1.26
Boom Truck		1		0.34
Scissor Lift		1		0.42
Face Charger, Explosives		2		0.84
Pneumatic Cartridge Loader		1		0.01
Pneumatic ANFO Loader		1		0.01
Blasting Utility		1		0.08
Shotcrete Sprayer		1		0.46
Personnel Carrier		3		0.24
Transmixer		1		0.48
Motor Grader		1		0.40
Utility Platform Lift		1		0.47
Fork Lift		1		0.11

Total		34		27.74

The Underground mine fixed plant includes primary and secondary ventilation fans, heaters, emergency egress system, dewatering equipment (pumps, piping, couplings, brackets, and support), workshop, electrical infrastructure (power distribution systems, cabling, controls and instrumentation, utilities, switches, and compressors), backfill system (grout tank and pump, grizzly, and rock box), and mine rescue equipment. Fixed plant costs are summarized in Table 21-18. Fixed plant sustaining costs are estimated at 10% of the initial purchase costs.

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Table 21-18: Underground Equipment Purchases—Fixed Plant


Fixed Plant Equipment	Total Cost ($M)
Ventilation and Egress		6.53
Dewatering Equipment		3.00
Electrical Infrastructure		7.49
Workshop, Services and Mine Rescue		2.06
Backfill System		0.66

Total		19.76

Pre-production period operating costs are capitalized during 2017 and 2018 and include mining labour, maintenance, power, heating, non-capital stope development, technical services, and operational planning.

Underground mine construction costs include civil, structural, mechanical and electrical installations, contract labour, and owner costs. These are grouped into the following cost items and are summarized in Table 21-19.

•

Underground Infrastructure and Foundations – including alimak support structure, portal substation, and fresh and return air raise foundations;

•

Underground Workshop, CAF Backfill Station and Ancillary Installations – including structural costs of the maintenance bay, sump bays, backfill station support, fuel bay, cranes and hoists;

•

Portal Entry and Infrastructure – including portal building foundations, excavation, backfill, and ground support; and

•

Owners Costs – including detailed design engineering, consulting engineering, field engineering, and site closure.

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Table 21-19: Contracted Construction Costs


Contracted Construction Costs	Total Cost ($M)
Underground Infrastructure and Foundations		2.01
Underground Workshop, CAF Backfill and Ancillary Installations		5.40
Portal Entry and Infrastructure		0.64
Owners Costs		4.99

Total		13.03

Project indirect costs are estimated at $9.7M and include freight, capital spares, commissioning, contractor supervision and training, and mobilization and demobilization.

21.5.6

Tailings Water Management and Treatment

Approximately $40.5M in sustaining capital is allocated for the tailings dams, water management infrastructure and a water treatment plant.

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.

The estimated cost to design, construct, install and commission the facilities for the water treatment plant (AMEC 2013) is $12.1M. This amount covers the direct field costs of executing the project, plus the indirect costs associated with design, construction and commissioning. Escalation of material and equipment costs has not been included in the estimate; however, a $1.6M contingency was applied. Mechanical equipment such as reactors, agitators, in-line lime silo, dosing systems, acid storage tank, a process water pumphouse and other minor equipment are included in the estimate. A 12 x 16 m building is required for the control room, electrical room and for housing of certain pieces of mechanical equipment such as the reagent pumps. Direct costs also include 323 m³ of concrete, 1,500 m of piping and other various costs pertaining to automation, civil/architectural, electrical and structural/miscellaneous steel. The cost for construction of a settling pond dam inside the Water Management Pond is also included.

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21.5.7

Fish Habitat Compensation Costs

Fish habitat compensation costs were provided by AMEC and total $1.99M. This cost includes onsite fish habitat for diversion channels and ponds and offsite watershed restoration works (livestock fencing and offline water).

21.5.8

Chapple Township Compensation Costs

Chapple Township compensation costs were provided by New Gold and total $2.1M. This cost includes Official Planning Amendment/Zoning By-Law Amendment (“OPA/ZBLA”), water pipeline easement compensation and 10.1 km of road closure compensation.

21.5.9

Salvage Value

Total equipment salvage value was calculated to be $60.5M and is credited partially during Year 10 (2026) once the open pit has finished and in Year 14 (2030) at the end of the Project. An estimate salvage value percentage was applied to the initial equipment value.

Year 10

After the open pit is depleted, open pit mining equipment not required for stockpile reclamation will be sold for a total estimated salvage value of $30.2M.

Year 14

At the end of the mine life, remaining equipment will be sold for a total estimated salvage value of $30.3M. The major equipment include the SAG and ball mill (including motors), primary crusher, pebble crusher and apron feeders along with the balance of the plant equipment.

21.5.10

Reclamation and Closure Costs

Reclamation costs are estimated to be $20.5M and closure costs are estimated to be $29.4M for a total of $49.9M, provided by AMEC. Closure and reclamation costs include monitoring inspection program, engineering, contracts, supervision, reporting, removal of 230 kV transmission line and poles, drainage works, site ponds, stockpiles reclamation, buildings, remote services, tanks, roads, etc.

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21.6

Operating Costs

21.6.1

Introduction

Operating costs were calculated based on open pit and underground mining, processing, refining, transport and general and administrative costs. Operating costs and any revenue from production incurred during the pre-production period until commercial production is achieved have been capitalized within New Gold’s Owner’s costs. A summary of the overall Project operating costs per year for each area is shown in Figure 21-1, Table 21-20 and Table 21-21.

Figure 21-1: Project Operating Costs

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Table 21-20: Annual Project Operating Cost Summary ($/t milled)


	2017		2018		2019		2020		2021		2022		2023		2024		2025		2026		2027		2028		2029		2030		TOT
Open Pit Mine		12.35		14.32		16.23		16.63		17.18		16.91		12.77		9.15		3.65		1.21		1.35		1.38		1.29		1.30		9.22
Underground Mine		0.00		0.00		4.16		4.09		5.62		6.00		6.05		5.81		5.74		6.17		4.26		1.33		0.00		0.00		3.63
Process Plant		9.13		9.27		9.27		9.07		9.16		9.33		9.24		9.33		9.26		9.24		9.36		9.28		9.21		9.38		9.25
General & Administration		1.65		1.62		1.60		1.60		1.60		1.58		1.58		1.58		1.57		1.51		1.51		1.36		1.35		1.38		1.54
Refining & Transport		0.14		0.14		0.18		0.17		0.20		0.19		0.16		0.15		0.13		0.09		0.10		0.10		0.08		0.09		0.14
Royalties		0.07		0.07		0.20		0.21		0.12		0.28		0.26		0.53		2.39		0.77		0.22		0.09		0.07		0.58		0.41

TOTAL		23.34		25.43		31.65		31.77		33.88		34.29		30.07		26.56		22.74		18.99		16.82		13.53		12.00		12.73		24.19

Table 21-21: Annual Project Operating Cost Summary ($M)


	2017		2018		2019		2020		2021		2022		2023		2024		2025		2026		2027		2028		2029		2030		TOT
Open Pit Mine		93.1		109.8		124.4		127.4		131.7		129.6		97.9		70.1		28.0		9.3		10.4		10.5		9.9		5.7		957.8
Underground Mine		0.0		0.0		31.9		31.3		43.1		46.0		46.4		44.5		44.0		47.3		32.7		10.1		0.0		0.0		377.2
Process Plant		68.8		71.1		71.1		69.5		70.2		71.5		70.9		71.6		71.0		70.8		71.7		70.9		70.6		41.4		961.0
General & Administration		12.5		12.4		12.3		12.3		12.3		12.1		12.1		12.1		12.1		11.6		11.5		10.4		10.4		6.1		160.0
Refining & Transport		1.1		1.1		1.4		1.3		1.5		1.5		1.2		1.2		1.0		0.7		0.8		0.7		0.6		0.4		14.4
Royalties		0.5		0.6		1.6		1.6		0.9		2.1		2.0		4.1		18.3		5.9		1.7		0.7		0.5		2.6		43.1

TOTAL		175.9		194.9		242.6		243.5		259.7		262.9		230.5		203.5		174.3		145.5		128.8		103.4		91.9		56.1		2,513.6

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Table 21-22 provides a summary of the average cash costs per tonne milled and ounce of gold over the LOM. Operating cash costs, including silver credits and royalties for the LOM total USD $663/oz. Au. All in sustaining cash costs, including silver credits and royalties for the LOM total USD $765/oz. Au.

Table 21-22: Project Cash Cost Summary


Area	Year 1 – 9 Cash Cost (USD/oz.)		LOM Cash Cost (USD/oz.)
Open Pit Mining (Waste + Rock +		390		373
Overburden + Stockpile Rehandling)¹		390		373
Underground Mining (Cut & Fill)²
Processing		207		268
General & Administration		36		45
Refining and Transportation		4		4
Royalty Payments		10		12

Subtotal Costs		646		702

Silver by-Product Sales		-33		-39

Total Costs Net Silver		613		663

Sustaining Capital		123		102

All-in Sustaining Cash Costs		736		765

^1.	Equivalent to $2.04/t mined, (excluding stockpile rehandling costs of $1.34/t mined).

^2.	Equivalent to $90.10/t mined.

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, due to the processing of stockpile material, the overall operating costs per ounce increase substantially during the later years with gold grades ranging from 0.3 to 0.6 g/t Au.

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

21.6.2

Power and Fuel

The cost of electrical power for the Project was determined to be $0.065/kWh. This includes a reduction for the Northern Industrial Rebate Program and is a projected 2014 rate based on historical rates from 2006 to 2013. The cost of propane and diesel, $0.50/L and $0.95/L, respectively was provided by local suppliers and used for cost calculations.

21.6.3

Total Employees

The number of employees during the production period (Years 1-14) consists of personnel from the open pit mine, underground mine, process plant and general and administrative. While the process plant and general and administrative employees remain constant through the mine life, the open pit and underground mine employees vary per year, as shown in Figure 21-3.

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Figure 21-3: Total Personnel

The total personnel for the Rainy River Project peaks in Year 5 (2021) at 606 employees as shown in Table 21-23.

Table 21-23: Project Peak Personnel


Area	Project Peak (Year 5)
Open Pit Mine		311
Underground Mine		172
Process Plant		91
General and Administrative		29

Total		606

21.6.4

Open Pit Operating Costs

Mine operating costs were estimated by BBA using the equipment list and manpower requirements based on the mine plan presented in Chapter 16. The mining operating costs include all costs related to ore extraction, waste, and overburden material and rehandling of the stockpiled material. All open pit mine production costs incurred during the pre-production phase of the Project will be capitalized. The open pit mining production period is from Year 1 to Year 9. Stockpile rehandling occurs from Year 9 until the end of the mine life (Year 14).

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Table 21-24 presents the total open pit mining cost summary for pre-production, production, and stockpile rehandling periods.

Table 21-24: Open-Pit Mining Cost Summary Per Period


	Units	Pre-Production (Years -2 and -1)	Production (Year 1 to 9)	Pre-Production and Production (Year -2 to Year 9)	Stockpile Rehandling (Years 9 to 14)
Mining Cost for
Rock (Ore, Waste)	$/t mined	2.11	2.11	2.11
Mining Cost for Overburden	$/t mined	1.54	1.52	1.53
Total Mining Cost	$/t mined	1.81	2.04	2.02
Rehandling Cost	$/t mined				1.34

The open-pit mine operating cost is presented by cost area in Table 21-25. The average operating cost over the LOM is $2.04 per tonne of material mined, including pre-production capitalized costs or $2.02 per tonne of material mined excluding pre-production capitalized costs.

Table 21-25: Open Pit Mine Operating Cost Breakdown


Cost Area	LOM Operating Cost ($M)		Operating Cost ($/t mined)
Equipment Maintenance	$	328.5	$	0.67
Equipment Fuel and Electricity	$	269.7	$	0.55
Blasting	$	152.5	$	0.31
Personnel	$	238.7	$	0.49
Services	$	5.0	$	0.01

Total Open Pit Operating Cost¹	$	994.4	$	2.02

Stockpile Rehandling	$	50.4	$	1.34

^1.	Total Mine Opex includes capitalized pre-production costs. Mine Opex during production years is $2.04/t mined and $1.81/t mined during pre-production years.

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Equipment Maintenance

Equipment operating costs were developed from vendor quotations. Other sources of information include an internal database of similar projects. It includes maintenance (e.g. repairs, spare parts, bucket repairs), chains, ground engaging tools grease and oil, tires, truck liners, drilling tools and accessories, etc. The total cost was compiled based on a cost per hour of operation for each type of equipment. Equipment maintenance costs exclude the cost of maintenance personnel, fuel, and electricity as they were calculated separately. Equipment maintenance over the LOM is $328.5M or $0.67/t milled.

Equipment Fuel and Electricity

Diesel fuel is used to operate mine trucks, hydraulic diesel shovels, loaders, dozers, and other smaller mine equipment. Yearly fuel consumption was estimated according to the annual operating hours, based on equipment specifications and truck haulage profiles. Electrical power is supplied to the open pit by a power distribution loop and is used to operate the electric hydraulic shovel. Yearly power consumption was estimated according to the annual operating hours, equipment specifications, and equipment utilization. Equipment fuel and electricity is $269.7M over the LOM or $0.55/t milled.

Blasting

Blasting costs for ore and waste rock have been estimated based on parameters and powder factors presented in Chapter 16. Blasting costs were estimated based on the blast pattern for required fragmentation and annual tonnage in both ore and waste. The blasting unit costs were derived from vendor quotations.

Emulsion costs were estimated at $0.29/t for ore and $0.23/t for waste rock, based on a unit cost of $82.20 per 100 kg of emulsion. It is assumed that 20% of the waste will be drilled using the ore pattern. Blasting costs also include accessories and contractor labour costs for mixing, delivering explosives to the blast holes, and loading explosives into the blast holes. Pre-splitting blasting costs are considered in the estimation of total blasting costs. Total blasting costs over the LOM are $152.5M or $0.31/t milled.

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Personnel

Labour requirements have been estimated on an annual basis to support the mine plan developed in this study. Mine salaried and hourly personnel positions with corresponding headcounts were presented in Chapter 16. Salaried and hourly personnel base salaries were provided by New Gold and included: a short-term incentive bonus, a long term incentive bonus, employment insurance, Canada pension plan, medical, workplace safety and insurance board and company pension fund factors. The combined burdens for the salaried personnel varied by position, averaging 31.5%. Total personnel costs including burden is $238.7M or $0.49/t milled.

Services

Service costs include items such as consultant fees, software licences, an allowance for mine dewatering, dispatch system licence, and survey and monitoring equipment. Services represent $5M over the LOM or $0.01/t milled.

Stockpile Re-handling

After depletion of the Open Pit, the mill will be fed at 21,000 tpd using the low-grade stockpiled material. The operating cost estimate for rehandling is $50.4M or $1.34/tonne of reclaimed ore material ($0.50/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.

21.6.5

Underground Operating Costs

The total operating cost for the underground mine is estimated to be $377M, or $90.10 per tonne of ore mined (average), and includes the costs of all key mine operating activities from the onset of production (2017) to the end of the mine life (2026).

The estimation of underground mine operating costs is based on a first principles work-up of all key mining activities, using the following inputs:

•

Mine design and method;

•

Production and development schedule;

•

Power and ventilation requirements; and

•

Waste and backfill schedule.

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The operating costs are divided into key mining activities as shown in Table 21-26.

Table 21-26: LOM Underground Operating Costs by Activity


Operating Costs	Total LOM Cost ($M)		Average LOM Cost ($ / tonne mined)
Waste Development		28.9		6.90
Ore Development		35.9		8.57
LH Stoping		48.8		11.65
Backfill		27.2		6.49
Mine Maintenance		14.7		3.51
Mine General		57.0		13.61
Labour		164.8		39.36

Total		377.2		90.10

Associated cost areas include equipment operating, explosive and fuel consumption, consumables (ground support and backfill) and services requirement (ventilation, cabling, and piping).

Equipment selection and operating costs are based on mine activities and workups of required operating hours, productivity and cycle times, availability/utilization, and fuel consumption. The operating costs for all major mobile equipment were obtained from vendor quotes and benchmarked against AMC’s cost database for similar recent projects. Ancillary equipment utilization and costs are calculated based on projected usage to complete general mine tasks.

The unit costs of all major consumables, including explosives, ground support, pipes, and ventilation ducting, have been obtained from vendor quotes. The unit costs of power, propane, cement, and diesel were provided by New Gold.

Mine Maintenance costs include fixed plant and electrical maintenance, maintenance overheads, and shop consumables. The Mine General and Mine Maintenance costs are developed based on industry experience and benchmarked against similar projects in the region.

Manpower requirements and labour rates, including benefits and burdens, were developed in consultation with New Gold. The manpower build-up is estimated based on mine activities and

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mobile equipment quantities. All underground mine operating activities during the first three (3) years are proposed to be completed by a contracted workforce. Labour rates in these years include a 22% contractor premium. After the initial three (3) years, all underground mine operating activities will be undertaken by New Gold and hence, no contractor premiums are applied after production begins.

Mine General

Mine General costs include power consuming items include primary and secondary ventilation fans, mobile drilling equipment, dewatering pumps, workshop services, and the underground CAF plant. Propane heating costs are based on the scheduled total mine airflow which increases annually in tandem with development and production activity until steady state production, 1,500 tpd, is achieved. The breakdown of Mine General costs is presented in Table 21-27.

Table 21-27: Underground Operating Costs – Mine General Area


Mine General	Total LOM Cost ($M)		Average LOM Cost ($ / tonne mined)
Power		23.2		5.53
Heating		13.0		3.11
Support Equipment		6.4		1.52
Technical Service Consumables		1.3		0.30
Mine General Consumables		2.1		0.50
Personnel Safety Equipment		0.6		0.15
Freight		1.8		0.42
Definition Drilling		8.7		2.08

Total		57.0		13.61

21.6.6

Process Plant

Process plant operating costs were calculated for 14 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 LOM processing operating costs were determined to be $9.25/t milled at approximately 21,000 tpd. The yearly tonnages from the mine plan vary and were used to obtain the operating costs.

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Operating costs during ramp-up/commissioning (up to 60% of 21,000 tpd production) were included in the capital expenditures as Owner’s Costs (Section 21.4.9).

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. A breakdown of the average processing operating costs can be seen in Table 21-28.

Table 21-28: Processing Operating Cost Breakdown


Cost Area	LOM Operating Cost ($M)		Operating Cost ($/t milled)
Reagents		187.4		1.80
Spare Parts/Maintenance		65.3		0.63
Liners and Screen Components		49.0		0.48
Grinding Media		155.7		1.50
Personnel		139.3		1.34
Electrical Power		288.3		2.77
Propane		10.6		0.10
Transportation (reagents, media)		57.7		0.56
Water Treatment		7.7		0.07

Total		961.0		9.25

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.

Reagents

The reagent consumptions were estimated based on testwork, industrial references, literature and assumed operational practice. The individual reagent costs ($/t reagent) were established through vendor quotations and comparison with prices at reference sites. Reagents represent approximately 20% of the total process operating cost at $187.4M or 1.80 $/t milled.

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Spare Parts / Maintenance

Vendors provided yearly quotes for various spare parts and the estimated required maintenance. Spare parts / maintenance represent approximately 7% of the total process operating cost at $65.3M or $0.63/t milled.

Liners and Screen Components

Liner and screen components were calculated based on vendor quotes for the estimated parts cost and their expected wear life, provided in months. Liners and screen components represent approximately 5% of the total process operating cost at $49.0M or $0.48/t milled.

Grinding Media

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 individual media costs ($/tonne steel media) were established through vendor quotations and comparison with prices at reference sites. Grinding media represents approximately 16% of the total process operating cost at $155.7M or $1.50/t milled.

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 New Gold. It was assumed that contractors would be used for crusher and liner changes.

There are a total of 91 employees accounted for in the process operating costs, 78 employees in the process plant and 13 employees in the assay lab.

The personnel costs represent approximately 15% of the total process operating cost at $139.3M or $1.34/t milled.

Electrical Power

The respective power required for the SAG and ball mill were calculated based on the material hardness (A x b value) for the SAG mill and the BWi of the ball mill. The power consumption of all the major equipment was calculated using the electrical load list and mechanical equipment list.

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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 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 A x b 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 2,800 µm to 75 µm. The A x b 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.

The electrical power costs represents approximately 30% of the total process operating costs at $288.3M or $2.77/t milled.

Propane

The propane consumption includes requirements for the crusher, concentrator, process and boiler. The propane costs represent approximately 1% of the total process operating costs at $10.6M or $0.10/t milled.

Transportation (Reagents, Media)

Transportation costs were provided by vendor quotes and displayed separately in the process operating cost to allow a better comparison of reagent and grinding media prices. The transportation costs represent approximately 6% of the total process operating costs at $57.7M or $0.56/t milled.

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Water Treatment Plant

An annual cost of $734k was calculated by AMEC for operating the water treatment plant starting in Year 5 until the end of the mine life. The water treatment plant costs represent approximately 1% of the process operating costs at $7.7M or $0.07/t milled.

21.6.7

G&A 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.

The G&A costs were calculated for 14 years of operation and with an average cost of $1.54/tonne milled. This cost includes:

•

Human Resources;

•

Site Administration, Insurance and Management;

•

Infrastructure Power;

•

Health and Safety Supplies;

•

First Nations Participation Agreement Payments;

•

Security and Paramedic Services;

•

Environmental Costs;

•

G&A Personnel;

•

Information Technology (“IT”); and

•

Training.

The labour costs include the twenty-nine (29) employees required for G&A, including nineteen (19) staff members and ten (10) hourly employees. In general, the management and administrative staff will work 40 hours per week on day shift. Warehousing personnel will work a 12-hour shift per day to support the 24 hours of required daily operations. The labour cost represents 25% of the G&A costs. Infrastructure electricity accounts for approximately 21% and site administration, maintenance and insurance represent approximately 28% of the G&A costs. The breakdown is shown in Table 21-29.

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Table 21-29: Average General and Administrative Costs


Cost Area	LOM Operating Cost ($M)		Operating Cost ($/t milled)
Infrastructure Electricity		34.1		0.33
Site Admin., Insurance and Maintenance		45.1		0.43
Health and Safety		7.5		0.07
Environment		6.9		0.07
Human Resources		11.0		0.11
IT and Telecommunications		10.2		0.10
Personnel		39.5		0.38
First Nations Participation Agreement Payments		5.7		0.06

Total		160.0		1.54

21.6.8

Royalties

The annual royalty costs are based on the Feasibility Study mine design and production profile, along with the terms of the individual royalty agreements for each property, which in turn relate to a limited portion of the reserves. Royalty costs are based on a gold price of USD $1,300 per ounce of gold and USD $22 per ounce of silver. Over the life of the Project, approximately $43M in royalties is expected to be paid.

21.6.9

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. New Gold will be paid for a set recovery of the assayed content. Over the LOM, a transport and refining cost of $14.4M or $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

22.1

Introduction

A financial analysis for the Rainy River 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 the total investment was calculated based on the assumption that no debt financing is required for the project. 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 performed for the pre-tax base case to assess the impact of variations of the project capital costs, annual operating costs, price of gold/silver and USD/CAD exchange rate.

The results of the economic analysis summarized below represent forward-looking information as defined under Canadian securities law. Actual results may differ materially from those expressed or implied by forward-looking information. The reader should refer to the Cautionary Note with respect to Forward Looking Information at the front of this Report for more information regarding forward-looking statements, including material assumptions (in addition to those discussed in this section and elsewhere in this Report) and risks, uncertainties and other factors that could cause actual results to differ materially from those expressed or implied in this section (and elsewhere in the Report).

22.2

Methods, Assumptions and Basis

The Economic Analysis was performed using the following assumptions and basis:

•

The Project NPV was determined on a pre-tax and after-tax basis with discounting to the start of Year -2 (2015), which marks the first year of project construction. Project expenses incurred in Year -3 (2014) are treated in the cashflow model as occurring at the start of Year -2;

•

Cashflows are assumed to occur mid-period;

•

The Project Executive Schedule developed in the Feasibility Study (Chapter 24) is used and takes into consideration key project milestones;

•

The Financial Analysis was performed for the Open Pit and Underground Proven and Probable Mineral Reserves estimated in this Study;

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•

Commercial production start-up is scheduled to begin in the first quarter (Q1) of 2017. Operations are estimated to span over a period of approximately 14 years;

•

The base case gold and silver prices are USD $1,300/oz. and USD $22/oz., respectively;

•

The United States to Canadian dollar exchange rate has been assumed to be USD $0.95:CAD $1.00 during preproduction and operation years;

•

All cost and sales estimates are in constant Q4 2013 Canadian dollars, with no inflation or escalation factors taken into account;

•

All gold and silver is sold in the same year it is produced;

•

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

•

Details of capital and operating costs are provided in Chapter 21 “Capital and Operating Costs”;

•

All values shown are post payment of royalties. Total royalty payments over the LOM are $28.6M NPI/NSR (net profit interest / net smelter return) and $14.5M other;

•

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 in the cashflow from New Gold’s operating mines;

•

Final rehabilitation and closure costs will be incurred after production Year 14;

•

After-tax results were provided by New Gold. BBA has not verified this work;

•

After-tax figures assume a combined income tax rate of 25% with 2.7% corporate minimum tax, a mining tax of 10% of taxable mining profits over $500,000 and an allocation of corporate tax attributes among New Gold’s operations; and

•

All dollars are in Canadian, unless specifically noted.

The general assumptions used for this financial model are summarized in Table 22-1, Table 22-2 and Table 22-3.

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Table 22-1: Financial Model Criteria and Production Summary^1,2


Area	Units	LOM Annual Average		Year 1-9 Annual Average		LOM Total
Gold Price	USD/oz		1,300		1,300		1,300
Silver Price	USD/oz		22		22		22
Ore Milled (OP & UG)	Mt		7.67		7.65		104.28
Open Pit Waste Mined	Mt		23.44		32.96		318.18
Strip Ratio (Waste: Ore)	—		3.18		3.00		3.18
Blended Gold Grade	g/t		1.12		1.44		1.12
Blended Silver Grade	g/t		2.80		3.07		2.80
Gold Recoveries	%		90.61		91.89		90.61
Silver Recoveries	%		64.05		64.15		64.05
Gold Recovered	koz		251		325		3,402
Silver Recovered	koz		442		480		6,004
Exchange Rate	CAD:USD		0.95		0.95		0.95
Discount Rate	%		5		5		5
Initial Capital Cost	$ M		—		—		931.58
Sustaining Capital	$ M		26.99		41.90		366.34
Life of Mine	years		—		—		13.6

^1.	Year 1 starts from commercial production (excluding pre-production).

^2.	Table excludes 0.4 Mt of material milled in the pre-production period.

Table 22-2: Operating Costs and Cash Costs over the LOM


Area	LOM Total ($M)		$ per tonne milled (LOM)		USD per gold ounce produced (LOM)
Open Pit Mining (Waste + Rock + Stockpile)		958		9.22		373
Under Ground Mining (Cut & Fill)		377		3.63		373
Processing		961		9.25		268
General & Administration		160		1.54		45
Refining and Transportation		14		0.14		4
Royalty Payments		43		0.41		12

Subtotal Costs		2,514		24.19		702

Silver by-Product Sales		-139		-1.34		-39

Total Costs Net Silver		2,375		22.86		663

Sustaining Capital		366		3.53		102

All-in Sustaining Cash Costs		2,741		26.38		765

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Table 22-3: Capital Costs over the LOM (M $CAD)


Description	Capital Costs ($M)
Initial Direct Costs		672
Initial Indirects		186
Initial Contingency		73

Total Pre-Production Capital Costs		931

Sustaining Capital		366

Total LOM Capital Costs		1,297

22.3

Royalties

The annual royalty costs were calculated by New Gold and are based on the conceptual Open Pit and Underground mine design and production profiles developed by BBA and AMC, 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$1,300 per ounce of gold and USD$22 per ounce of silver. Over the life of the Project, approximately $43.1M in royalties is expected to be paid based on the base case metal prices and project assumptions, with $28.6M NPI/NSR (net profit interest / net smelter return, respectively) and $14.5M in other royalties.

22.4

Salvage Value

Total equipment salvage value was calculated to be $60.5M and is credited partially during Year 10 (2026) once the open pit is depleted, and in Year 14 (2030), at the end of the Project. In Year 10, the open pit mining equipment not being used for stockpile reclamation will be sold for a total estimated salvage value of $30.2M. In year 14, at the end of the mine life, the remaining mine and process plant equipment will be sold for a total estimated salvage value of $30.3M. Major equipment includes the SAG and ball mills (including motors), primary crusher, pebble crusher and apron feeders, along with the balance of the plant equipment.

22.5

Taxation

The tax rate calculations and taxation assumptions used in the after-tax NPV calculations are summarized below.

The federal government imposes income tax on mining income at the same rate that applies to other types of income. The federal rate applicable to resource profits is 15%. Ontario’s taxation

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of the resource sector is generally harmonized to the federal system. The provincial corporate income tax rate applicable to mining income is 10%. A combined rate of 25% is used in the model to compute the federal and provincial tax liability in respect of the Rainy River Project. In addition, in the earlier years of the project, the Ontario Corporate Minimum Tax has been computed at a rate of 2.7%. All deductions and rates are based on currently enacted legislation. In addition, Ontario’s Mining Tax Act is levied at a rate of 10% on annual taxable profits in excess of $500,000.

The federal and provincial tax legislation provides a number of deductions, allowances, and credits that are specifically available to taxpayers engaged in qualifying mining activities. The most notable of these deductions are Canadian Exploration Expenditures (“CEE”), Canadian Development Expenses (“CDE”), and capital costs eligible for Class 41 of the capital cost allowance system. Because these deductions and allowances are only available when incurred, a high-level assumption was made with regard to the allocation of expenditures between the three categories in the life-of-mine model.

Similarly, the Ontario Mining Tax Act provides a number of deductions in arriving at taxable profits, the key ones being allowance for Exploration and Development Expenditures, Depreciation Allowance and Processing allowance. Since these deductions depend on when the expenses are incurred and the degree of processing that occurs in Canada, a number of high level assumptions are made with regards to the allocation and timing of expenditures for the depreciation allowance calculation. A general assumption regarding the processing allowance has been made and it has been assumed that Rainy River will qualify for a minimum 15% processing allowance.

Rainy River Resources Limited is a 100% subsidiary of New Gold Inc. (“New Gold”). New Gold is intending to implement tax planning strategies that will allow it access to Rainy River’s tax attributes, with the overall goal of maximizing the overall profitability of New Gold’s Canadian operations as a whole, as opposed to any one operation or project. Part of New Gold’s growth strategy, including the successful acquisition of Rainy River Resources in 2013, has been to build its business in jurisdictions where it already has an established presence. One of the many benefits of this approach is that it enables the company to manage its business in a tax-efficient manner. In the case of Rainy River, New Gold plans to implement tax planning strategies that

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would allow it to realize tax synergies by utilizing a portion of the tax attributes that have been, and will continue to be generated in the corporate entity holding the company’s portfolio of Canadian assets. Currently, the company’s primary objective is to maximize the cash flow generation of its New Afton Mine in Kamloops, British Columbia, with any remaining tax attributes planned to be used to maximize Rainy River’s future after-tax cash flow. As New Gold’s focus in such allocation will be to maximize the company’s overall profitability rather than that of any one operation or project, this will remain a dynamic process.

At the end of 2012, on a combined basis, the New Gold corporate entity had over $900 million in Canadian Exploration Expenditures, Canadian Development Expenditures and Class 41 capital cost allowance, as well as over $100 million in net operating losses. These totals exclude those attributes specifically related to the Blackwater project. In addition, the company’s annual corporate administration expenses are approximately $30 million and its annual interest expenses are $52 million, both of which can be combined with the above noted attributes to shelter taxable profits from one or both of New Afton and Rainy River going forward. New Gold intends to implement tax planning strategies that would allow it the flexibility to access the existing tax attributes of Rainy River Resources and utilize them in a manner which would maximize New Gold’s overall profitability.

22.6

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 22-4.

Table 22-4: Financial Analysis Summary (Pre-tax and After-tax)


Description		Base Case		Units
	Net Present Value (0% disc)		983		$M
	Net Present Value (5% disc)		462		$M
	Internal Rate of Return		13.1			%
	Simple Payback Period¹		5.4		Years
	Net Present Value (0% disc)		774		$M
	Net Present Value (5% disc)		330		$M
	Internal Rate of Return		11.3			%
	Simple Payback Period¹		5.5		Years

^1.	Payback period starts from commercial production using undiscounted cash flows.

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The pre-tax base case financial model resulted in an internal rate of return of 13.1%, a NPV of $986M with a discount rate of 0%, and a NPV of $462M with a discount rate of 5%. The simple payback period is 5.4 years. On an after-tax basis, the base case financial model resulted in an internal rate of return of 11.3% and a NPV of $774M with a discount rate of 0%, and a NPV of $330M with a discount rate of 5%. The simple payback period is 5.5 years. Figure 22-1 shows the gold production, cash costs and the all in cash costs. The all-in cash costs in Years 10 (2026) and 14 (2030) include a credit for the estimated equipment salvage values. The Year 10 salvage credit consists of the sale of all mine equipment not required for the stockpile reclamation operation. The Year 14 salvage credit consists of the sale of all remaining mine and process plant equipment at the end of the life of mine.

Figure 22-1: Life-of-Mine Cash Flow Projection

The summary of the Rainy River Project discounted cash flow financial model (pre-tax and after tax) is presented in Table 22-5.

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Table 22-5: Rainy River Financial Model Summary (CAD$)

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Figure 22-2 shows the Project’s cumulative cash flows projected for the mine life on a pre-tax and after-tax basis.

Figure 22-2: Life-of-Mine Cash Flow Projection (Pre-tax and After-tax, discount rate: 5%)

22.7

Sensitivity Analysis

A financial sensitivity analysis was conducted on the Project’s base case cash flow NPV and IRR for variations in capital expenditures, operational expenditures, metal recoveries and metal prices. Table 22-6 shows the pre-tax and after-tax NPV (in CAD$) and IRR sensitivity to varying metal prices and exchange rates. The results are also presented in USD$ in Table 22-7. The royalty payments were assumed to be constant (equivalent to the base case) for the sensitivity analysis.

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Table 22-6: Sensitivity Results for Metal Price and Exchange Rate Variations (CAD $M)


Gold Price (USD$ / ounce)		Silver Price (USD$ / ounce)		Exchange Rate (USD$:CAD$)		NPV at 5% Discount Rate (CAD$ 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
	1,150		20		0.93		147		107		7.8		7.1		6.8		6.8
	1,300		22		0.95		462		330		13.1		11.3		5.4		5.5
	1,450		24		0.97		763		536		17.6		14.9		4.3		4.4
	1,600		26		1.00		1,013		706		21.1		17.8		3.6		3.8

Table 22-7: Sensitivity Results for Metal Price and Exchange Rate Variations (USD $M)


Gold Price (USD$ / ounce)		Silver Price (USD$ / ounce)		Exchange Rate (USD$:CAD$)		NPV at 5% Discount Rate (USD$ 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
	1,150		20		0.93		138		100		7.8		7.1		6.8		6.8
	1,300		22		0.95		438		314		13.1		11.3		5.4		5.5
	1,450		24		0.97		738		520		17.6		14.9		4.3		4.4
	1,600		26		1.00		1,009		706		21.1		17.8		3.6		3.8

The pre-tax sensitivity analysis for Project capital costs, operating costs and metal recovery is summarized in the following table.

Table 22-8: Selected Sensitivities, Pre-Tax NPV at 5% Discount Rate (CAD $M)


	Sensitivities
Description	-30%			-20%			-10%		0%		10%		20%			30%
LOM Capital Expenditures		818			699			581		462		343		225			106
LOM Operating Expenditures		975			804			633		462		291		120			(51	)
Gold Price		(525	)		(196	)		133		462		791		1120			1449
Silver Price		434			443			453		462		471		481			490
USD Foreign Exchange Rate		1477			1139			800		462		124		(215	)		(553	)
Open Pit Mining Costs		742			649			555		462		369		275			182
Underground Mining Costs		594			550			506		462		418		374			330
Processing Costs		656			591			527		462		397		333			268

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The results of the financial sensitivity analysis are graphed in Figure 22-3. It can be seen that the Project financials are most sensitive to the gold price and foreign exchange rate.

Figure 22-3: Sensitivity of the Net Present Value (Pre-tax) to Selected Financial Variables

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23.

ADJACENT PROPERTIES

23.1

Introduction

Seven (7) properties in the exploration stage are located adjacent to or near the Rainy River Project property. Bayfield Ventures Corp. (TSX-V:BYV) holds three (3) of the properties, known as the B Block, C Block and Burns Block. Coventry Resources Inc. (TSX-V:CYY) holds three (3) of the properties, known as the Pattullo, Nelles and Blue properties. King’s Bay Gold Corp. (TSX-V:KBG) holds the seventh property, being a single continuous land package contiguous with the most northerly portion of the Rainy River Project property. The closest Canadian operating mines are the Lac des Iles; palladium, nickel, gold and copper mine, owned by North American Palladium Ltd. (TSX:PDL) and Goldcorp’s (TSX:G) Red Lake mine complex. Both projects are approximately 400 km from the Rainy River Project.

The following information was obtained from public sources, and neither New Gold nor BBA has verified its accuracy. The information provided in the following section is not a reflection of the mineralization of the Rainy River Project property.

Bayfield Ventures Corp.

Bayfield Ventures Corp. (“Bayfield”) is exploring for gold and silver in the Rainy River district and owns a 100% interest in mineral rights to three (3) properties adjacent to the Rainy River Project property. These consist of the B Block, C Block and the Burns Block.

The Burns Block is an 80 acres parcel of land directly east of the ODM Zone and west of the newly discovered Intrepid Zone, which is completely surrounded by properties either owned by, or optioned to New Gold (both surface and mineral rights). New Gold also owns a 100% interest in the surface rights to the Burns Block.

On January 14^th, 2014, Bayfield issued a press release: “Bayfield Announces Maiden NI 43-101 Technical Report and Mineral Resource Estimate on Burns Block Rainy River Gold-Silver Project, NW Ontario”. The press release stated combined Indicated mineral resources from both potential open pit and underground mineralization to be 60 koz. Au and 686 koz. Ag and 151 koz. and 1,563 koz Ag of Inferred resources. An NI 43-101 compliant technical report for the Burns Block entitled “Burns Block National Instrument 43-101 Compliant Technical Report” was filed on SEDAR on January 23, 2014.

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Bayfield also published a NI 43-101 technical report in March 2007 on its Tait Gold project located 3 km southwest of the Rainy River project 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.

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 Project, known as the Pattullo, Nelles and Blue properties.

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).

King’s Bay Gold Corp.

King’s Bay Gold Corp. (“King’s Bay”) is exploring for gold in the Kenora Mining District in Northern Ontario. King’s Bay owns a 100% interest (both surface and mineral rights) to the Menary gold projects representing 18 claims (2,238 hectares) approximately 20 km southeast of the town of Nestor Falls, Ontario and approximately 15 km northeast of the Rainy River project.

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24.

OTHER RELEVANT DATA AND INFORMATION

24.1

Project Execution

The Project Execution Strategy was developed for the Rainy River Project Feasibility Study (FS) based on the latest FS information available and industry 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 New Gold’s 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 New Gold’s 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 2015.

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. 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

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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. New Gold plans to operate the Project in accordance with the International Cyanide Management Code.

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.

Remaining basic engineering and detailed engineering begins early in the Project. Construction work packages will be issued on a fixed-price basis dependent on the 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, project management team, 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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Figure 24-1: Project Management Organization Chart

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24.6

EPCM Project Controls and Reporting

Project Controls will be implemented to provide the project management and the Owner with timely, updated reports on the project status for analysis and control, ensure that the project scope is completed on time and within budget including; budgeting, incurred cost, actual cumulative capital cost, final cost forecasts, reasons for variances, and monthly cash flows. The periodically updated information will include updates of scope, budget, commitments, schedule, and trends.

Baseline Definition

At the start of basic definition phase, the EPCM contractor will assemble a task force of experienced project personnel who will complete documentation for the “Baseline Definition”. This documentation will include:

•

Project execution plan;

•

Project schedule;

•

Project cost estimate;

•

Project control budget;

•

List of engineering deliverables;

•

Project Manpower Forecasting and Leveling (“MFL”);

•

Project Manual including project instructions.

Scope and Engineering Management

Once the Baseline Definition is completed, the project enters the execution phase and any variations will be documented through coordination meetings and various types of reviews. Engineering Deliverables will be progressed on a monthly basis, earned value and forecast hours to complete calculated.

Cost and Change Management

Once the budget structure and the baseline schedule have been completed and approved by the project management team and the Owner, the baseline budget will be developed and serve as reference for project cost control and management. Change management will include tracking all changes through the use of trends and project change notices.

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Document Management

The EPCM contractor will put in place a Document Management system that will consist of controlling the document reception and distribution for tender, purchase, construction and review.

Project Status Report

Monthly and weekly reporting will be established as a standard tool to track and monitor progress and budget.

24.7

Project Scheduling

Along with the FS, the Project will produce a detailed project schedule that 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 22-month project construction duration assumes commencement of field activities in January 2015 and mechanical completion in November 2016. The detailed engineering is scheduled to start in January 2014 to allow sufficient progress to award fixed-price construction contracts. The purchase of major process equipment is assumed to be completed in January 2014 to allow the detailed design of buildings and infrastructures.

The January start schedule specifically takes into account 2014/2015 winter work for activities such as bulk earthworks and concrete placement that 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 January 2015. Detailed engineering is expected to achieve substantial completion in the second quarter (Q2) of 2015, which will facilitate the Project Team’s ability to maximize Lump Sum Construction Contracts.

The key Project milestones are summarized in Figure 24-2.

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Figure 24-2: Project Milestones

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After acceptance of the FS and the associated trade-off studies, the implementation of the Project Schedule will begin by ordering outstanding long-lead capital equipment items critical to engineering design. At the same time, detail engineering and remaining basic engineering for the plant and infrastructure will start.

24.8

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.

24.9

Site Development

Highway 600

Construction of Highway 600 will be the subject of a specific road contract. The East Access Road will be included in the major civil contract.

230 kV Overhead Line

The construction of the 230 kV power line will be contracted to a specialized contractor on a design-build basis.

24.10

Construction

24.10.1

Construction Management Responsibilities

The CM group will be responsible for the management of the construction site under the authority of the EPCM Project Manager. The Construction Manager will be responsible for effectively planning, organizing, and managing the construction quality, safety, budget, and scheduling objectives of the Project.

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.

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24.10.2

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 mid-2015. It is anticipated that 230 kV power will be available in July 2015. Pre-commissioning and commissioning of electrical distribution power lines and equipment will commence thereafter. From August 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.

24.10.3

Construction Labour Requirement

The RRP construction 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 1,670,000 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 447 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 2015-2016. The slowdown reflects the project schedule objective, while reducing the expenses related to winter concrete work.

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Figure 24-3: Monthly Construction Manpower Graph

24.11

Process Facilities

24.11.1

Critical Path and Installation Methodology

The Schedule has been presented in association with the established Project Work Breakdown Structure (“WBS”) that 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 delivered to the crusher area via the mine trucks. This work will be completed by the fourth quarter (Q4) of 2015.

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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.

24.12

Tailings Management Area (“TMA”) 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 that 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.

24.13

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.

During the latter part of engineering, the Owner and the CM/Pre-Commissioning team will develop a commissioning plan. 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 Pre-Commissioning

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team for dry and wet commissioning. The Pre-Commissioning team will consist of EPCM staff, plant operators and maintenance staff and 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 for the commencement of ramp-up. The transfer of systems will be formally documented and will include all mechanical/electrical testing documents and supplier information.

24.14

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 lockout is ensured, hazards identified, and communication protocols established.

24.15

Risk Management

The formal risk management program began during the initial 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 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 Project can support a 19,500 tpd open pit and a 1,500 tpd underground mine. Support is based on the total direct processing of Proven and Probable Mineral Reserves of 66.9 Mt grading 1.54 g/t Au and 3.26 g/t Ag, and 37.4 Mt of stockpile Proven and Probable Mineral Reserves grading 0.38 g/t Au and 1.99 g/t Ag. The QP’s have read the following conclusions and made the following interpretations:

25.1

Sampling Method, Approach and Analyses

Rainy River used industry best practices to collect, handle and assay core samples collected during the period from 2005 to 2013. 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

The 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.

Upon completion of the validation procedures, SRK concludes that the digital database for the Rainy River 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 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 1,665 core boreholes (742,424 m), which includes 230 boreholes (79,575 m) drilled within the Intrepid Zone during the period from 1996 to August 16, 2013, which has been added since the previous October 2012 mineral resource model. The mineral resource statement effective November 2, 2013 is shown in Table 25-1.

SRK considers the mineral resource model document herein to be sufficiently reliable to support engineering and design studies to evaluate the viability of a mining project at a feasibility level.

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

SRK Consulting, November 2, 2013^1,2,3,4,5


	Quantity ‘000t		Grade				Metal
Category	Quantity ‘000t		Au gpt		Ag gpt		Au ‘000 oz		Ag ‘000 oz
Direct Processing Material
Open Pit²
Measured		20,282		1.45		1.93		947		1,261
Indicated		80,411		1.35		2.55		3,486		6,584
O/P Measured & Indicated		100,693		1.37		2.42		4,433		7,846
Inferred		9,388		0.97		2.28		292		687
Underground²
Measured		89		4.95		2.75		14		8
Indicated		5,469		4.53		11.34		796		1,994
Measured & Indicated		5,558		4.53		11.20		810		2,002
Inferred		2,641		4.46		8.30		379		707
Stockpile Material³
Open Pit
Measured		6,294		0.37		1.29		74		262
Indicated		64,816		0.44		2.17		919		4,526
Measured & Indicated		71,110		0.43		2.09		993		4,788
Inferred		8,626		0.37		1.16		102		323
Combined Direct Processing and Stockpile Mineral Resources
Measured		26,665		1.21		1.79		1,035		1,531
Indicated		150,696		1.07		2.70		5,202		13,104
Measured and Indicated		177,361		1.09		2.57		6,236		14,635
Inferred		20,655		1.16		2.58		773		1,717

¹	Mineral resources are reported in relation to conceptual pit shells which are limited to 150m below sea level and are inclusive of the Intrepid Zone.

³	Direct processing material is defined as mineralization above a cut-off of 0.45 g/t gold and likely to be mined and processed directly.

⁴

⁵	Qualified Persons – The mineral resource statement was prepared by Dorota El-Rassi, P.Eng. (APEO #100012348) and Glen Cole (APGO #1416) from SRK Consulting (Canada) Inc., both Independent “Qualified Persons” as that term is defined in Canadian National Instrument 43-101.

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A comparison between the October 2012 and the November 2013 Mineral Resource Statements is shown in Table 25-2. The reduction in Measured and Indicated Open Pit Mineral Resources grade is primarily due to the increase in block size in 2013 and the reduction in reporting cut-off grade. The reduction in Inferred resources is due to a difference in mineral resource reporting methodology in relation to a conceptual pit shell. The increase in the Underground Mineral Resources is primarily due to the addition of the Intrepid Zone to the Mineral Resource Statement.

Table 25-2: Comparison of October 2012 and November 2013 Mineral Resource Statements


	Quantity			Grade (g/t)						Contained Metal (oz.)
Classification	(tonnes)			Gold			Silver			Gold			Silver
Open Pit
Measured		-4	%		-10	%		-6	%		-13	%		-9	%
Indicated		+15	%		-12	%		-11	%		0	%		+3	%
Measured & Indicated		+11	%		-13	%		-9	%		-2	%		1	%
Inferred		-81	%		-6	%		-22	%		-82	%		-85	%
Underground
Measured		+1	%		+0	%		+0	%		+0	%		+0	%
Indicated		+32	%		+1	%		+85	%		+33	%		+144	%
Measured & Indicated		+31	%		+1	%		+85	%		+32	%		+143	%
Inferred		+194	%		+7	%		+79	%		+216	%		+428	%

25.4

Sampling Preparation, Analysis and Security

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 grade 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: Open Pit and Underground Proven and Probable Mineral Reserves^1,2,3,4,5,6


Mineral Reserves Category	Tonnage (‘000 t)		Au Grade (g/t)		Ag Grade (g/t)		Contained Metal Au (koz)		Contained Metal Ag (koz)
Direct Processing Material
Open Pit
Proven		15,839		1.47		2.04		746		1,038
Probable		46,866		1.26		3.05		1,896		4,594
Underground
Proven		—		—		—		—		—
Probable		4,187		4.96		10.31		668		1,388
Total Direct Processing Material		66,892		1.54		3.26		3,311		7,021
Stockpile Material
Open Pit
Proven		6,843		0.38		1.51		84		332
Probable		30,541		0.39		2.10		378		2,058
Total Stockpile Material		37,384		0.38		1.99		462		2,390
Combined Direct Processing and Stockpile Material
Open Pit
Proven		22,681		1.14		1.88		830		1,370
Probable		77,407		0.91		2.67		2,275		6,652
Total		100,088		0.96		2.49		3,105		8,022
Underground
Proven		—		—		—		—		—
Probable		4,187		4.96		10.31		668		1,388
Total		4,187		4.96		10.31		668		1,388
Total Combined
Proven		22,681		1.14		1.88		830		1,370
Probable		81,594		1.12		3.06		2,943		8,040
TOTAL		104,275		1.13		2.81		3,773		9,410

^1.

Open pit mineral reserves have been estimated using an optimized pit shell based on metal prices of USD $800 per ounce gold and USD $25 per ounce silver, a foreign exchange rate of CAD $1.05 to USD $1.00, gold recovery of 89.9% (non-CAP zone) and 74.3% (CAP zone) and a silver recovery of 67.1% (non-CAP zone) and 69.5% (CAP zone). The cut-off grade is based on a gold price of USD $1,200. Underground reserves have been estimated from mining shapes generated using a cut-off grade of 3.5 g/t gold-equivalent. Development material from stope access drives above a cut-off grade of 1.5 g/t gold-equivalent is also assumed to be sent to the mill for processing. Underground breakeven cut-off grade calculated at 2.75 g/t gold-equivalent based on metal prices of USD $1,300 per ounce gold and USD $22 per ounce silver, a foreign exchange rate of CAD $1.05 to USD $1.00, gold recovery of 95% and a silver recovery of 75%.

^2.

^3.	Open pit direct processing material is defined as mineralization likely to be mined and processed directly and above a variable cut-off grade ranging from 0.3-0.7 g/t.

^4.	Stockpile material includes all material within designed open pit between variable cut-offs described above in Note 3, as well as material within the CAP zone (code 500) that is suitable for stockpiling and future processing.

^5.

Mineral Reserves for the open pit are derived from the resource model effective November 2, 2013. Models for the underground reserves were derived from the August 2013 and September 2013 models for the main ODM zone and Intrepid Zone, respectively. Models were 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 Canadian National Instrument 43-101. The combined mineral resource statement, including the Intrepid Zone, was provided to BBA on November 2, 2013. Rainy River’s exploration program in Richardson Township is being supervised by Mark A. Petersen, (AIPG Certified Professional Geologist #10563), Vice President, Exploration for New Gold and a “Qualified Person” as defined in Canadian National Instrument 43-101. New Gold continues to implement a rigorous QA/QC program to ensure best practices in drill core sampling, analysis and data management.

^6.

Qualified persons -The open pit portion of the mineral reserve statement was prepared under the supervision of Patrice Live (OIQ #38991) of BBA, and the underground portion of the mineral reserve statement was prepared by Colm Keogh, P.Eng. (APEGBC #37433) of AMC Mining Consultants (Canada) Ltd., both independent “Qualified Persons” as that term is defined in Canadian National Instrument 43-101.

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Open Pit Mining

Conventional open pit mining has been chosen as the primary method to extract the Rainy River deposit because of the deposit’s geometry and proximity to the surface. The primary equipment fleet at the peak of operations consists of: three (3) hydraulic shovels (2 x 26 m³, 1 x 29 m³), one (1) wheel loader (18 m³), 22 haul trucks (220 t class), three (3) DTH drills (8.5”) and a fleet of support equipment. Operating bench heights of 10 m have been planned for mining operations. Over the life of the mine, a total of 318.2 Mt of waste rock and 73.6 Mt of overburden will be moved. The operational open pit stripping ratio, excluding waste and overburden stripping during the development phase is 3.5:1.0. It is anticipated that all overburden will be removed within the first seven (7) years of mine operation. Mined waste and overburden will be stored in nearby stockpiles or used in dam construction activities associated with the Tailings Management Area (TMA) or blended in cement and used as backfill material for the underground mine.

The LOM mining operating cost is estimated at $2.04/t mined (including waste, ore and overburden). Open pit mining operating costs include 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 the Feasibility Study and from BBA’s recently updated database. At the peak of open pit mining, a workforce of 318 persons (staff and hourly employees) will be required.

Underground Mining

A main decline from a surface portal located to the east of the open pit will be used to access the mine. A fleet of 7 m³Load Haul Dump trucks (“LHDs”) and 45-tonne trucks will be used for material loading and transport from the various underground working areas through an internal ramp system that connects all levels to the main decline. Loading will occur in close proximity to the stoping areas and ore will be hauled directly to a surface coarse ore stockpile adjacent to the portal.

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An extensive underground development program that attains a peak of 560 m/month of jumbo advance is required to develop and maintain access to adequate resources to sustain 1,500 tpd of ore production. The ramp-up period to full production will require five (5) years from the onset of mine development. Waste generated through infrastructure development will be disposed of in underground stopes whenever operationally practical, however, an estimated excess of 1.80 Mt must be hauled to the surface waste stockpile over the life of mine.

The LOM average underground operating cost has been estimated to be $90.10 /t ore mined. Underground mining costs include labour, mine general costs, LH stoping, ore development, waste development, backfill and mine maintenance. A maximum of 176 people will be required in Year 5 (2021) and 168 people in the Project peak year (Year 5, 2021).

25.6

Metallurgy, Processing and General & Administrative

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 original 2013 Feasibility Study and for this 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 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.

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The design A x b 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 and consultant used to size the grinding circuit provides a good level of confidence in the sizing of the SAG and ball mills. 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 0.6% (average), ranging from 90-91% (LOM: 90.6%), and a silver recovery including solution losses of approximately 2-3%, ranging from 63-65% over the life of mine (LOM: 64.1%). It is anticipated that, over a mine life of 14 years, approximately 3,402 koz. of gold and 6,004 koz. of silver will be produced.

The average life-of-mine processing operating costs were determined to be $9.25 per tonne milled. The operating costs are based on metallurgical testwork, the mine plan, New Gold salary compensation/benefit guidelines, and recent supplier quotations. The operating costs include reagents, consumables, grinding media, personnel (including contractors), electrical power, and propane. A full personnel list has been developed and it is expected that 91 people will be required for the processing plant. This number is not expected to vary significantly over the life-of-mine.

The average life-of-mine General and Administrative (“G&A”) operating costs were determined to be $1.54 per tonne milled. The operating costs include Human Resources, Site Administration, Management and Insurance, Infrastructure Power, Health and Safety Supplies, First Nations Participation Agreement Payments, Security and Paramedic Services, Environmental Costs, Personnel, Information Technology (“IT”) and Training. A full personnel list has been developed and it is expected that 29 people will be required for G&A. This number is not expected to vary significantly over the life-of-mine.

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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 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.

Tailings Management and Dam Design

The total volume of tailings produced over the mine life will be approximately 75 Mm³at 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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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.

Electricity

The total power demand of the Project was determined to be approximately 58 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

The initial capital cost of the Project is estimated to be $931M; with a sustaining capital of $366M. The life-of-mine all-in sustaining cash cost is USD $765/oz. Au, including silver credits and royalty payments. The life-of-mine operating cash cost is USD $663/oz. The pre-tax Project NPV is estimated to be $462M using a discount rate of 5% and an after-tax NPV of $330M using a discount rate of 5%. The Project’s pre-tax internal rate of return (IRR) is estimated to be 13.1% and the pre-tax payback period of the Project is 5.4 years.

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New Gold compiled the taxation calculations for the Project. The federal rate applicable to resource profits is 15%. Ontario’s taxation of the resource sector is generally harmonized to the federal system. The provincial corporate income tax rate applicable to mining income is 10%. A combined rate of 25% is used in the model to compute the federal and provincial tax liability in respect of the Rainy River Project. As well, in the earlier years of the project, the Ontario Corporate Minimum Tax was computed at a rate of 2.7%. All deductions and rates are based on currently enacted legislation. In addition, Ontario’s Mining Tax Act is levied at a rate of 10% on annual taxable profits in excess of $500,000.

After tax NPV is estimated to be $330M using a discount rate of 5%. The Project IRR (after tax) is estimated to be 11.3% and the simple after tax payback period is 5.5 years.

The pre-tax discounted sensitivity analysis indicates that positive project returns can be achieved over the likely range of variation in capital costs (± 30%), operating costs (-30% / +27%), gold prices (-14% / +30%), and a USD foreign exchange rate (-14% / +30%). 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 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 Chapter 26. However, the final decision to proceed with a construction of a mining operation at the Rainy River Project site is at the discretion of New Gold.

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26.

RECOMMENDATIONS

•

Continuation of open pit and underground mine design optimization;

•

Consideration of an underground mining test program within the Intrepid Zone;

•

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;

•

Continuation of key personnel recruitment;

•

Continuation of coordinated environmental assessment process;

•

Continuation of First Nation, Métis and public consultations;

•

Continuation of hydrogeological studies in specific areas;

•

Secure required permits and authorizations from government and regulatory agencies;

•

Continuation of value and detailed engineering activities; and

•

Construction of the Project (following appropriate approvals).

26.1

Proposed 2014 Work Program Budget

In order to advance the Project according to the proposed schedule and this feasibility study, the 2014 budgeted costs for initiation of key project activities are estimated at approximately US $104 M, as shown in Table 26-1.

Table 26-1: Budget for 2014 (US dollars)


Activity	Cost (US $ M)
Capitalized exploration and condemnation drilling		14.0
Property, plant and equipment payments		60.0
Detailed engineering and studies		20.0
Permitting and environmental monitoring		10.0

Total		104.0

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Feasibility Study of the Rainy River Project

As of the effective date of this report, New Gold has already budgeted, authorized and initiated a considerable amount of the recommended work as stated in their press release dated February 6, 2014 “New Gold Finishes 2013 with Lowest Costs in its History, Increase Gold Reserves by 127 Percent per Share and Provides 2014 Guidance with Even Lower Costs”

BBA has reviewed New Gold’s 2014 proposal for detailed engineering work, permitting and exploration activities on the Rainy River Project property and considers that the program budget is reasonable. BBA recommends that New Gold implement the program as in order to maintain the proposed Project schedule as described within this Feasibility Study. It should be noted that initiation of these proposed activities and the overall budget are subject to funding, permitting or other matters that may cause the proposed program to be altered in the normal course of New Gold’s business activities, or alterations which may affect the program as a result of the activities themselves. The reader should refer to the Cautionary Note with respect to Forward Looking Information at the front of this Report for more information regarding forward-looking statements, including material assumptions (in addition to those discussed in this section and elsewhere in this Report) and risks, uncertainties and other factors that could cause actual results to differ materially from those expressed or implied in this section (and elsewhere in the Report).

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2014 Detailed 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 (February 2014). Further development of the Rainy River Project property should address key risks and opportunities that have been identified through the Feasibility Study work program.

Geology and Exploration

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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;

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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;

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SRK supports and understands New Gold’s wish to consolidate and re-interpret currently available data with the objective to optimize data extraction and knowledge to help guide future local exploration and regional exploration efforts: and

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Condemnation drilling to support mine infrastructure design.

New Gold has budgeted approximately US $14 M for exploration related activities to test the potential of the company’s land package in 2014. The program is scheduled to include 35,000 to 40,000 metres of exploration drilling of which approximately 70% will be dedicated to testing new targets around known ore bodies and along newly recognized district trends. The remaining 30% of the drilling is for completion of condemnation drilling within the immediate project development area. This proposed condemnation/exploration drilling program is considered to be aligned with the recommendations of this Study.

Open Pit and Underground Mining

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To review the current overburden material stockpiling approach;

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Revisit the elevated open pit cut-off grade strategy using the current mill throughput and mining method to maximize the positive NPV impact;

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Continue with value engineering initiatives;

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To review the open pit slope as more data is collected and available;

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To develop a detailed operational mine plan for pre-production and for the first two (2) years of operation;

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Consider an underground bulk sample program within the Intrepid Zone to validate ground conditions, operating assumptions, grade distribution, and to reduce the level of uncertainty associated with the future larger scale investment in mine development. Consideration should be given to commencing this program earlier than the current scheduled start of underground activity in 2017;

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Further optimization of the underground mine design and schedule should be undertaken following the test mining program; and

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Underground cut-off grade optimization work should be undertaken in tandem with the mine design optimization work.

AMC estimates that $30M would be sufficient funds for the bulk sample program work that includes:

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Portal entry construction.

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2900 m of lateral development under contract, of which nearly 500 m would be within mineralized horizons. A 100 m waste drift to permit definition drilling has also been provided.

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208 m of raise development for ventilation and secondary egress.

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All stationary equipment including fans, dewatering equipment, electrical distribution and refuges.

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5000 m of definition drilling.

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Owner’s costs and a 15% contingency have been included in the $30M estimate. It should be noted that costs for the work described above have been included in the mine-wide underground development capital estimate (Section 21).

Open Pit Geotechnical Design

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 east to west and north to south with either:

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Additional televiewer surveys or;

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Two (2) additional boreholes either side of the pit centre towards the north.

Underground Geotechnical Design

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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;

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All further exploration cores 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;

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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.

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Geotechnical investigation should be performed for the portal location and for the first 300 m of the ramp;

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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 and backfill raise;

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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; and

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It is recommended that once the ramp has reached a depth approaching 500 m, that in-situ stress over coring tests are performed preferably at two locations separated vertically by 100 m to 150m.

Metallurgical Testing and Processing

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Hydraulic testing for the final confirmation of tailings pumping characteristics and thickener settling characteristics;

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Continue with value engineering initiatives;

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Additional testwork for thickener sizing and flocculant dosages to optimize sizing of thickeners and flocculant system;

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Additional testwork for equipment sizing, as required;

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Investigate the design circulating load for the ball mill circuit to potentially improve grinding efficiency.

Infrastructure

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Continue with value engineering initiatives;

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Prepare a site material balance by month to confirm usage and source of material for roads, backfill and aggregates; complete the assessment of available quantity of material from the former MOT borrow pit;

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Submit application for electrical power to IESO Ontario.

Environmental and Tailings Management

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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;

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Geochemical characterization should continue as the processing and mining plans are detailed, with modification to the mineral waste management plan as appropriate;

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Estimated reclamation costs and bonding requirements should be reassessed in the next phase of development as more detailed engineering designs become available;

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The design of the water treatment plant at the TMA should be optimized during detailed design (Section 18.9.2);

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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 local clay borrow areas and the mine waste overburden suitable for dam construction

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27.

REFERENCES

AMEC, 2011. Rainy River Project, 2011 Wildlife Baseline Study.

AMEC, 2012. Estimated Groundwater Seepage into Proposed Underground Mine Workings at the Rainy River Gold Project. Technical Memorandum for Rainy River Resources Ltd. Submitted November 15, 2012.

AMEC, 2012A. Rainy River Gold Project - Site Investigations. Report - Rock Mechanics Site Investigation/Joint Mapping and Rock Mass Characteristics (TC113921) version Draft. Document prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd. on May 4, 2012.

AMEC, 2012a. Overburden Material Delineation, Rainy River Gold Project - Feasibility Study. Technical Memorandum TC121506 prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd., November 2, 2012.

AMEC, 2012B. Rainy River Gold Project - Site Investigations. Preliminary Summary – Preliminary Rock Mechanics Overview (TC113921) Version 1. Document prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd. on July 27, 2012.

AMEC, 2012C and 2012c. Rainy River Gold Project - Site Investigations. Report - 2011 Geotechnical and Hydrogeological Investigation Report (TC113921) Vversion 3. Document prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd. on August 8, 2012.

AMEC, 2012D. Rainy River Resources Limited - Rainy River Feasibility Study Rock Mechanics Site Investigation Laboratory Testing Of Intact Rock Core (Draft - TC113921) Document Prepared by AMEC Environment & Infrastructure. Submitted to Rainy River Resources Ltd. on August 17, 2012.

AMEC, 2012E. Rainy River Gold Project - Feasibility Study. Drawing - Pit Design Zone Recommendations -Pit Shell (TC121506). Document prepared by AMEC Environment & Infrastructure, submitted to BBA on October 3, 2012.

AMEC, 2012F and 2012f. Rainy River Gold Project - Feasibility Study. Open Pit Overburden Slope Design Considerations (TC121506) Vversion 3. Document prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd. on October 31, 2012.

AMEC, 2012G. 2012 Hydrogeological Baseline Study (TC111504), Rainy River Gold Project, Rainy River Resources Limited, Draft report, prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd. on December 17, 2012.

AMEC, 2012H. Rainy River Aquatic Resources 2011 Baseline Investigation.

AMEC. 2012I. Rainy River Project, 2011 Wildlife Baseline Study.

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AMEC. 2012J. Rainy River Project, 2011 Species at Risk Report.

AMEC. 2012K. Rainy River Project, 2012 Terrestrial Baseline Study.

AMEC. 2012L. Rainy River Project, 2012 Species at Risk Report.

AMEC, 2013A. Rainy River Underground Ground Support Recommendations Memo (Draft - TC121506) Document prepared by AMEC Environment & Infrastructure. Submitted to Rainy River Resources Ltd. and Golder Associates Limited, March 2013.

AMEC, 2013a. Geotechnical and Hydrogeological Site Investigations, Rainy River Gold Project. Report TC113921.100 by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd., February 26, 2013.

AMEC, 2013B. Rainy River Cemented Rockfill Feasibility Study Test Program - Interim Report (Draft – TC121506). Document prepared by AMEC Environment & Infrastructure, Submitted to Rainy River Resources Ltd., April 2013.

AMEC, 2013b. Geotechnical Recommendations for Plant Site Foundations Rainy River Gold Project - Feasibility Study. Report TC121506.100.6 prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd., January 17, 2013.

AMEC, 2013C. Rainy River Resources Limited – Rainy River Feasibility Study Construction Aggregate Source Assessment (Draft – TC121506). Document prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd., April 2013.

AMEC, 2013c. Geotechnical Design of the Tailings, Mine Waste and Water Management Facilities – Rainy River Gold Project Feasibility Study. Draft Report TC121506.100.23 by AMEC Environment and Infrastructure, submitted to Rainy River Resources Ltd., April, 2013.

AMEC, 2013D. Rainy River Resources Limited - Rainy River Feasibility Study Rock Mechanics Site Investigation Laboratory Testing of Intact Rock Core (Final - TC113921) Document prepared by AMEC Environment & Infrastructure. Submitted to Rainy River Resources Ltd., April 2013.

AMEC, 2013d. Proposed clay test embankment program, Rainy River Gold Project - Feasibility Study. Technical Memorandum TC121506 prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd., January 12, 2013.

AMEC, 2013E. Rainy River Resources Limited - Rainy River Feasibility Rock Mass Characterization Summary Of Geomechanical Site Investigation Program 2012 (Final - TC113921). Document prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd., April 2013.

AMEC, 2013e. Water Management Plan – Rainy River Gold Project Feasibility Study. Report TC121506.100.33 by AMEC Environment and Infrastructure, submitted to Rainy River Resources Ltd., April, 2013.

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AMEC, 2013F. Rainy River Resources Limited - Rainy River Feasibility Rock Mechanics Open Pit Stability Analysis Report (Final – TC121506). Document prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd., May 2013.

AMEC, 2013G. Rainy River Resources Limited - Rainy River Feasibility Rock Mechanics Underground Mine Design Report (Final – TC121506). Document prepared by AMEC Environment & Infrastructure, submitted to Rainy River Resources Ltd., May 2013.

AMEC 2013h. Treatability Tests for Cd and Zn Removal Rainy River Gold, dated April 2013.

AMEC 2013i. Rainy River Gold Project Feasibility Study: Water Management Pond, dated May 2013.

AMEC 2013j. Rainy River Project – Feasibility Level Treatment Plant Design, in progress.

AMEC 2013k Clay borrow assessment in Water Management Pond area (Doc3098004—004000-A1-ETR-0001-AA)

AMEC, 2013l. Water Management Plan – Rainy River Gold Project Feasibility Study. Report by AMEC Environment and Infrastructure, submitted to Rainy River Resources Ltd., October, 2013.