EX-99.1 2 blackwatertech_2012.htm BLACKWATER TECHNICAL REPORT

blackwatertech_2012.htm

Exhibit 99.1

CERTIFICATE OF QUALIFIED PERSON

Herbert E. Welhener, MMSA - QPM

Independent Mining Consultants Inc.

3560 East Gas Road

Tucson, Arizona  85714  USA

Tel:  +1 (520) 294-9861

E-mail: hwelhener@imctucson.com

I, Herbert E. Welhener, MMSA - QPM, am employed as a principal mining engineer with Independent Mining Consultants Inc.

This certificate applies to the technical report titled “Blackwater Project, British Columbia, Canada, NI 43-101 Technical Report on Preliminary Economic Assessment” with an effective date of August 28, 2012 (the “technical report”).

I am a Qualified Professional Member (Mining and Ore Reserves) of the Mining and Metallurgical Society of America (#01307QP) and I am a Registered Member of the Society of Mining, Metallurgy, and Exploration, Inc. (# 3434330RM) both recognized as a professional association as defined by NI 43-101.  I graduated with the follow degree from the University of Arizona:  Bachelors of Science – Geology, 1973.

I have worked as a mining engineer or geologist for 37 years since my graduation from the University of Arizona.  I have been directly involved with mine design and scheduling, mine equipment selection and costing, and resource development and evaluation.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101).

I have not visited the Blackwater project site.

I am responsible for Sections 1.15.1, 1.15.3, 16.1, 16.2, 16.6 to 16.8, 21.1.2, 21.4.2, and 24.2.3 of the technical report.

I am independent of New Gold Inc. as independence is described by Section 1.5 of NI 43-101.

I have had no previous involvement with the Blackwater Project.

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

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 contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated:  October 9, 2012

“Signed and Sealed”

Herbert E. Welhener, MMSA - QPM

CERTIFICATE OF QUALIFIED PERSON

Bruno Borntraeger, P.Eng.

Knight Piésold Ltd.

750 W. Pender Street, Suite 1400

Vancouver, BC   V6C 2T8

Tel:  (604) 685-0543

E-mail:  bborntraeger@knightpiesold.com

I, Bruno Borntraeger, P. Eng., am employed as a Specialist Engineer/Project Manager with Knight Piésold Ltd.

This certificate applies to the technical report titled “Blackwater Project, British Columbia, Canada, NI 43-101 Technical Report on Preliminary Economic Assessment” with an effective date of August 28, 2012 (the “technical report”).

I am a Professional Engineer (20926) with the Association of Professional Engineers and Geoscientists of British Columbia.  I graduated with a Bachelor of Applied Science in Geological Engineering from the University of British Columbia, May 1990.

I have practiced my profession continuously for 22 years.  I have been directly involved in geotechnical engineering, mine waste and water management, mine development with practical experience from feasibility studies, detailed engineering, construction, operations and closure.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101).

I visited the Blackwater project sites on September 28, 2011; December 20, 2011; February 21, 2012; and April 24, 2012.

I am responsible for Sections 1.15.2, 1.17, 1.19, 1.21, 5.0, 9.6, 10.7, 16.3 to 16.5, 18.3 to 18.5, 18.9, 21.1.5, 24.1.3, and 24.2.2 of the technical report.

I am independent of New Gold Inc. as independence is described by Section 1.5 of NI 43–101.

I have had no previous involvement with the Blackwater Project.

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

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 contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated:  October 9, 2012

“Signed and Sealed”

Bruno. Borntraeger, P. Eng.

CERTIFICATE OF QUALIFIED PERSON

Ramon Mendoza Reyes, P.Eng.

AMEC Americas Limited

111 Dunsmuir Street, Suite 400

Vancouver, BC   V6B 5W3

Tel:  (604) 664-3075

E-mail: ramon.mendoza@amec.com

I, Ramon Mendoza Reyes, P.Eng., am employed as a Principal Mining Engineer with AMEC Americas Limited.

This certificate applies to the technical report titled “Blackwater Project, British Columbia, Canada, NI 43-101 Technical Report on Preliminary Economic Assessment” with an effective date of August 28, 2012 (the “technical report”).

I am a Professional Engineer in the province of British Columbia.  I graduated in 1989 from the National Autonomous University of Mexico with a bachelor’s degree in Mining Engineering, and in 2003 completed a M.Sc. Degree in Mining & Earth Systems Engineering from the Colorado School of Mines in Golden, Colorado, USA.

I have practiced my profession continuously for 23 years and have previously been involved with mine designs, mine planning, mine project economic evaluations and mine operations for precious metals, base metals, disseminated sulphide and industrial mineral projects in North America and South America.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101).

I did not visit the Blackwater project site.

I am responsible for Sections 1.1 to 1.3, 1.22 to 1.28, 2.0, 3.0, 4.9, 4.10, 19.0, 21.1.1, 21.1.3, 21.1.6 to 21.1.9, 21.2, 21.3, 21.3.1, 21.4.1, 21.4.4, 21.4.5, 21.5, 22.0, 24.1, 24.1.4, 24.2, 25.0, 26.0, and 27.0 of the technical report.

I am independent of New Gold Inc. as independence is described by Section 1.5 of NI 43–101.

I have had no previous involvement with the Blackwater Project.

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

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 contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated:  October 9, 2012

“Signed and Sealed”

Ramon Mendoza Reyes, P.Eng.

CERTIFICATE OF QUALIFIED PERSON

Ronald G. Simpson, P.Geo.

GeoSim Services Inc.

1975 Stephens St.

Vancouver, BC, Canada V6K 4M7

Tel:  (604) 803-7470

E-mail:  rgs@universe.com

I, Ronald G. Simpson, P.Geo., am employed as a Professional Geoscientist with GeoSim Services Inc.

This certificate applies to the technical report titled “Blackwater Project, British Columbia, Canada, NI 43-101 Technical Report on Preliminary Economic Assessment” with an effective date of August 28, 2012 (the “technical report”).

I am a Professional Geoscientist (19513) with the Association of Professional Engineers and Geoscientists of British Columbia.  I graduated with a Bachelor of Science in Geology from the University of British Columbia, May 1975.

I have practiced my profession continuously for 37 years.  I have been directly involved in mineral exploration, mine geology and resource estimation with practical experience from feasibility studies.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101).

I visited the Blackwater project sites on December 13, 2010; September 8, 2011; November 28, 2011; and September 20, 2012.

I am responsible for Sections 1.5 to 1.9, 1.10, 1.12 to 1.14, 4.1 to 4.5, 6.0, 7.0, 8.0, 9.1 to 9.5, 9.8, 9.9, 10.1 to 10.6, 10.9, 10.10, 11.1, 11.3 to 11.9, 12.0, 14.0, 15.0, 23.0, 24.1.1, and 24.2.1 of the technical report.

I am independent of New Gold Inc. as independence is described by Section 1.5 of NI 43–101.

I have had prior involvement with the property that is the subject of the Technical Report, the nature of which involves the preparation of a technical report prepared for Richfied Ventures Corp. and Silver Quest Resources Ltd. dated 2 March, 2011 and titled “Technical Report, Blackwater Gold Project, British Columbia, Canada”. The report was readdressed to New Gold Inc. on June 6, 2011. An updated Technical Report with the same title was completed for New Gold and Siver Quest on November 2, 2011.  An further update was completed for New Gold on March 23, 2012.

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

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 contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated:  October 9, 2012

“Signed and Sealed”

Ronald G. Simpson, P.Geo.

CERTIFICATE OF QUALIFIED PERSON

Ignacy (Tony) Lipiec, P.Eng.

AMEC Americas Limited

111 Dunsmuir Street, Suite 400

Vancouver, BC   V6B 5W3

Tel:  (604) 664-3130

E-mail: tony.lipiec@amec.com

I, Ignacy (Tony) Lipiec, P.Eng., am employed as a Process Manager with AMEC Americas Limited.

This certificate applies to the technical report titled “Blackwater Project, British Columbia, Canada, NI 43-101 Technical Report on Preliminary Economic Assessment” with an effective date of August 28, 2012 (the “technical report”).

I am a Professional Engineer in the province of British Columbia.  I graduated from the University of British Columbia with a B.A.Sc. degree in Mining & Mineral Process Engineering, in 1985.

I have practiced my profession for 27 years, and have been involved with metallurgical design and process engineering for gold, base metal and disseminated sulphide projects in North America and South America

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101).

I did not visit the Blackwater project site.

I am responsible for Sections 1.4, 1.11, 1.16, 1.18, 1.20, 4.6 to 4.10, 9.7, 10.8, 11.2, 13.0, 17.0, 18.1 to 18.2, 18.6 to 18.8, 18.10, 20.0, 21.1.4, 21.4.3, 24.1.2, and 24.2.4 of the technical report.

I am independent of New Gold Inc. as independence is described by Section 1.5 of NI 43–101.

I have had no previous involvement with the Blackwater Project.

I have read NI 43–101 and the sections of the technical report for which I am responsible have been prepared in compliance with that Instrument.

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 contain all scientific and technical information that is required to be disclosed to make those sections of the technical report not misleading.

Dated:  October 9, 2012

“Signed and Sealed”

Ignacy (Tony) Lipiec, P.Eng.

IMPORTANT NOTICE

This report was prepared as a National Instrument 43-101 Technical Report for New Gold Inc. (New Gold) by AMEC Americas Ltd. (AMEC).  The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in AMEC’s 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 contract with AMEC.  Except for the purposed 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.

New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

CONTENTS

1.0	SUMMARY			1-1
	1.1	Key Outcomes		1-1
	1.2	Location, Climate, and Access		1-2
	1.3	Mineral Tenure, Surface Rights, Royalties, and Agreements		1-2
	1.4	Environment, Permitting, and Socio-Economics		1-3
	1.5	Geology and Mineralization		1-5
	1.6	Exploration		1-6
	1.7	Exploration Potential		1-7
	1.8	Drilling		1-7
	1.9	Sample Analysis and Security		1-8
	1.10	Data Verification		1-10
	1.11	Metallurgical Testwork		1-11
	1.12	Mineral Resource Estimate		1-12
	1.13	Mineral Resource Statement		1-14
	1.14	Drilling Completed Since Database Closeout Date		1-16
	1.15	Proposed Mine Plan		1-16
		1.15.1	Pit Design and Scheduling	1-16
		1.15.2	Geotechnical Considerations	1-18
		1.15.3	Equipment	1-19
	1.16	Process Design		1-19
	1.17	Waste Management		1-20
	1.18	Infrastructure		1-21
	1.19	Water Management		1-22
	1.20	Power and Electrical		1-22
	1.21	Water Supply		1-22
	1.22	Capital Costs		1-23
	1.23	Operating Costs		1-24
	1.24	Markets		1-26
	1.25	Economic Analysis		1-26
	1.26	Sensitivity Analysis		1-30
	1.27	Interpretation and Conclusions		1-31
	1.28	Recommendations		1-31
2.0	INTRODUCTION			2-1
	2.1	Terms of Reference		2-1
	2.2	Qualified Persons		2-1
	2.3	Site Visits and Scope of Personal Inspection		2-1
	2.4	Effective Dates		2-3
	2.5	Information Sources and References		2-3
	2.6	Previous Technical Reports		2-3
3.0	RELIANCE ON OTHER EXPERTS			3-1
	3.1	Mineral Tenure		3-1
	3.2	Surface Rights		3-1
	3.3	Taxation		3-2
4.0	PROPERTY DESCRIPTION AND LOCATION			4-1

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

	4.1	Property and Title in British Columbia		4-1
		4.1.1	Mineral Title	4-1
	4.2	Tenure History		4-4
	4.3	Mineral Tenure		4-4
	4.4	Surface Rights		4-7
	4.5	Royalties and Encumbrances		4-9
		4.5.1	Davidson Claims	4-10
		4.5.2	Dave Option	4-10
		4.5.3	Jarrit Option	4-10
		4.5.4	JR Option	4-10
		4.5.5	Saulnier Claim	4-10
	4.6	Permits		4-11
	4.7	Environmental Liabilities		4-11
	4.8	Social License		4-11
	4.9	Significant Risk Factors		4-11
	4.10	Comments on Section 4		4-11
5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY			5-1
	5.1	Accessibility		5-1
	5.2	Climate		5-1
	5.3	Local Resources and Infrastructure		5-1
	5.4	Physiography		5-2
	5.5	Regional Seismicity		5-3
	5.6	Sufficiency of Surface Rights		5-3
	5.7	Comments on Section 5		5-3
6.0	HISTORY			6-1
7.0	GEOLOGICAL SETTING AND MINERALIZATION			7-1
	7.1	Regional Geology		7-1
	7.2	Project Geology		7-3
	7.3	Mineralization		7-5
	7.4	Comments on Section 7		7-8
8.0	DEPOSIT TYPES			8-1
	8.1	Comments on Section 8		8-5
9.0	EXPLORATION			9-1
	9.1	Grids and Surveys		9-1
	9.2	Geological Mapping		9-1
	9.3	Geochemical Sampling		9-1
	9.4	Geophysics		9-2
	9.5	Petrology, Mineralogy, and Research Studies		9-2
	9.6	Geotechnical and Hydrological Studies		9-2
	9.7	Metallurgical Studies		9-2
	9.8	Exploration Potential		9-2
	9.9	Comments on Section 9		9-4
10.0	DRILLING			10-1
	10.1	Legacy Drilling		10-1
	10.2	Drill Methods		10-6

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

	10.3	Geological Logging		10-6
	10.4	Recovery		10-7
	10.5	Collar Surveys		10-7
	10.6	Downhole Surveys		10-7
	10.7	Geotechnical and Hydrological Drilling		10-7
	10.8	Metallurgical Drilling		10-7
	10.9	Sample Length/True Thickness		10-8
	10.10	Comments on Section 10		10-9
11.0	SAMPLE PREPARATION, ANALYSES, AND SECURITY			11-1
	11.1	Sampling Methods		11-1
	11.2	Metallurgical Sampling		11-1
	11.3	Density Determinations		11-1
	11.4	Analytical and Test Laboratories		11-2
	11.5	Sample Preparation and Analysis		11-2
	11.6	Quality Assurance and Quality Control		11-3
	11.7	Databases		11-4
	11.8	Sample Security		11-4
	11.9	Comments on Section 11		11-4
12.0	DATA VERIFICATION			12-1
	12.1	Site Visit Verification		12-1
	12.2	QA/QC Review		12-1
		12.2.1	Standards	12-1
		12.2.2	Blank Samples	12-2
		12.2.3	Field Duplicates	12-2
		12.2.4	Coarse Reject Duplicates	12-3
		12.2.5	Pulp Duplicates	12-3
		12.2.6	Check Assays	12-3
	12.3	Database Verification		12-4
	12.4	Comments on Section 12		12-4
13.0	MINERAL PROCESSING AND METALLURGICAL TESTING			13-1
	13.1	Metallurgical Testwork		13-1
		13.1.1	Crushing and Grinding Testwork	13-2
		13.1.2	Mineralogy	13-3
		13.1.3	Gravity Testwork	13-4
		13.1.4	Whole Ore Leaching Testwork	13-4
		13.1.5	Flotation, Regrind, and Concentrate Leaching Testwork	13-6
		13.1.6	Heap Leach Testwork	13-9
	13.2	Process Alternatives Assessment		13-10
		13.2.1	Flowsheet Alternatives	13-10
	13.3	Recovery Estimates		13-12
		13.3.1	Recovery Estimate for Treatment of Low-Grade Mineralized Material	13-12
		13.3.2	High Grade Silver Mineralized Material	13-12
	13.4	Metallurgical Variability		13-13
	13.5	Deleterious Elements		13-13
	13.6	Comments on Section 13		13-13
14.0	MINERAL RESOURCE ESTIMATES			14-1
	14.1	Key Assumptions/Basis of Estimate		14-1
	14.2	Geological Models		14-1

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

	14.3	Exploratory Data Analysis		14-4
	14.4	Density Assignment		14-6
	14.5	Grade Capping/Outlier Restrictions		14-6
	14.6	Composites		14-7
	14.7	Variography		14-8
	14.8	Estimation/Interpolation Methods		14-8
	14.9	Block Model Validation		14-10
	14.10	Classification of Mineral Resources		14-10
	14.11	Reasonable Prospects of Economic Extraction		14-11
		14.11.1	Determination of Cut-off Grade	14-12
		14.11.2	Gold Equivalency	14-12
	14.12	Mineral Resource Statement		14-12
	14.13	Factors That May Affect the Mineral Resource Estimate		14-14
	14.14	Drilling Completed Since Database Closeout Date		14-14
	14.15	Comments on Section 14		14-15
15.0	MINERAL RESERVE ESTIMATES			15-1
16.0	MINING METHODS			16-1
	16.1	Pit Design		16-1
		16.1.1	Mine Plan	16-1
		16.1.2	Consideration of Net Smelter Return Values	16-2
		16.1.3	Pit Slope Angles	16-3
		16.1.4	Pit and Phase Design	16-4
		16.1.5	Pit Definition Floating Cone	16-4
		16.1.6	Phase Design	16-4
		16.1.7	Pit and Dump Maps	16-7
	16.2	Proposed Mine Production Schedule		16-7
	16.3	Waste Management		16-10
	16.4	Geotechnical Parameters		16-13
		16.4.1	Proposed Open Pit Area	16-14
		16.4.2	Proposed Tailings Storage Facility Areas	16-17
		16.4.3	Proposed Plant Site Area	16-17
		16.4.4	Aggregate Potential	16-18
	16.5	Hydrogeology		16-18
	16.6	Mining Equipment		16-18
		16.6.1	Operating Time per Shift	16-19
		16.6.2	Material Characteristics	16-22
	16.7	Mine Personnel		16-22
		16.7.1	Salaried Staff	16-22
		16.7.2	Hourly Personnel	16-22
	16.8	Comments on Section 16		16-25
17.0	RECOVERY METHODS			17-1
	17.1	Proposed Process Flow Sheet		17-1
	17.2	Plant Design		17-1
		17.2.1	Crushing and Coarse Ore Stockpile	17-3
		17.2.2	Grinding	17-4
		17.2.3	Gravity Separation	17-5
		17.2.4	Leaching and CIP	17-5
		17.2.5	Carbon Desorption and Regeneration and Gold Recovery	17-5

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

		17.2.6	Tailings Thickening	17-6
		17.2.7	Cyanide Destruction	17-6
	17.3	Energy, Water, and Process Materials Requirements		17-6
		17.3.1	Reagents	17-6
		17.3.2	Energy and Water	17-7
		17.3.3	Cyanide	17-7
	17.4	Comments on Section 17		17-8
18.0	PROJECT INFRASTRUCTURE			18-1
	18.1	Road and Logistics		18-1
	18.2	On-Site Project Infrastructure		18-1
	18.3	Waste Storage Facilities		18-6
		18.3.1	Waste Rock Classification	18-6
		18.3.2	Site Selection	18-8
	18.4	Proposed Tailings Storage Facilities		18-8
		18.4.1	Tailings Classification	18-8
		18.4.2	Site Selection	18-8
		18.4.3	Design Criteria	18-9
		18.4.4	Construction and Deposition Strategy	18-10
		18.4.5	Seepage Control Measures	18-12
		18.4.6	PAG Disposal Area	18-12
		18.4.7	NAG Disposal Area	18-13
		18.4.8	Closure and Reclamation Considerations	18-13
	18.5	Water Management		18-14
		18.5.1	Sequence of Construction	18-16
		18.5.2	TSF Water Management Plan	18-18
		18.5.3	Open Pit Water Management	18-18
		18.5.4	Reclamation and Closure	18-18
		18.5.5	Water Balance	18-19
		18.5.6	Seepage Collection and Recycle Ponds	18-19
		18.5.7	Water Reclaim System	18-20
	18.6	Camps and Accommodation		18-20
		18.6.1	Camp Facilities	18-20
		18.6.2	Employee Housing and Transportation	18-21
	18.7	Power and Electrical		18-21
	18.8	Fuel		18-23
	18.9	Water Supply		18-23
	18.10	Comments on Section 18		18-24
19.0	MARKET STUDIES AND CONTRACTS			19-1
	19.1	Market Studies		19-1
	19.2	Commodity Price Projections		19-1
	19.3	Contracts		19-1
	19.4	QP Conclusions and Comments on Section 19		19-1
20.0	ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT			20-1
	20.1	Baseline Studies		20-1
		20.1.1	Air Quality and Noise	20-1
		20.1.2	Terrestrial Environment	20-1
		20.1.3	Aquatic Habitat	20-2
		20.1.4	Archaeology	20-5

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

	20.2	Environmental Issues		20-5
	20.3	Environmental Liabilities		20-7
		20.3.1	Existing Environmental Liabilities	20-7
		20.3.2	Exploration Reclamation Bonding	20-8
		20.3.3	Other Environmental Liabilities	20-8
	20.4	Closure Plan		20-8
	20.5	Permitting		20-9
	20.6	Considerations of Social and Community Impacts		20-11
	20.7	First Nations Engagement		20-11
	20.8	Other Stakeholder Consultation		20-12
	20.9	Discussion on Risks to Mineral Resources and Mineral Reserves		20-12
	20.10	Comments on Section 20		20-13
21.0	CAPITAL AND OPERATING COSTS			21-1
	21.1	Initial Capital Cost Estimates		21-1
		21.1.1	Basis of Estimate	21-1
		21.1.2	Mine Capital Costs	21-3
		21.1.3	Infrastructure Capital Costs	21-3
		21.1.4	Process Capital Costs	21-5
		21.1.5	Waste Management Costs	21-5
		21.1.6	Road Access Capital Costs	21-6
		21.1.7	Owner’s Capital Costs	21-8
		21.1.8	Indirect Capital Costs	21-8
		21.1.9	Contingency	21-9
	21.2	Sustaining Capital		21-10
	21.3	Mine Closure Salvage Value		21-11
	21.4	Capital Cost Summary		21-11
	21.5	Operating Cost Estimates		21-12
		21.5.1	Basis of Estimate	21-12
		21.5.2	Mine Operating Costs	21-13
		21.5.3	Process Operating Costs	21-14
		21.5.4	General and Administrative Operating Costs	21-17
		21.5.5	Operating Cost Summary	21-18
	21.6	Comments on Section 21		21-19
22.0	ECONOMIC ANALYSIS			22-1
	22.1	Methodology Used		22-1
	22.2	Financial Model Parameters		22-2
		22.2.1	Assumptions	22-2
		22.2.2	Taxes	22-4
		22.2.3	Working Capital	22-4
	22.3	Pre-Tax Financial Results		22-5
	22.4	Post-Tax Financial Results		22-8
	22.5	Sensitivity Analysis		22-8
	22.6	Comments on Section 22		22-10
23.0	ADJACENT PROPERTIES			23-1
24.0	OTHER RELEVANT DATA AND INFORMATION			24-1
	24.1	Risks		24-1
		24.1.1	Conversion of Mineral Resources to Mineral Reserves	24-1
		24.1.2	Processing	24-1

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		24.1.3	Waste Management	24-1
		24.1.4	Project Economics	24-2
	24.2	Opportunities		24-2
		24.2.1	Delineation of Additional Mineralization	24-2
		24.2.2	Pit Slopes	24-2
		24.2.3	Mining	24-3
		24.2.4	Process Flowsheet and Parameters	24-3
25.0	INTERPRETATION AND CONCLUSIONS			25-1
26.0	RECOMMENDATIONS			26-1
	26.1	Phase 1 - Testwork		26-2
		26.1.1	Drill Programs	26-2
		26.1.2	Camp Operations	26-3
		26.1.3	Administration Costs	26-3
		26.1.4	PEA Engineering Studies	26-4
		26.1.5	Metallurgical Testwork	26-4
		26.1.6	Engineering Optimization Studies	26-4
	26.2	Phase 2 - Studies		26-4
		26.2.1	Feasibility Studies	26-4
		26.2.2	Environmental and Permitting	26-5
		26.2.3	Government and Community	26-5
		26.2.4	Project Development Planning	26-5
27.0	REFERENCES			27-1
	27.1	References		27-1
	27.2	Glossary		27-3
		27.2.1	Abbreviations	27-3
		27.2.2	Glossary Table	27-5
		27.2.3	Chemical Symbols Table	27-11
		27.2.4	Metric Measurements Table	27-13

Tables

Table 1-1:	Blackwater Project Indicated Mineral Resource Summary (Effective Date July 27, 2012, Ronald G. Simpson P.Geo)	1-15
Table 1-2:	Blackwater Project Inferred Mineral Resource Summary (Effective Date July 27, 2012, Ronald G. Simpson P.Geo)	1-15
Table 1-3:	Initial Capital Cost Summary by Major Area	1-24
Table 1-4:	Summary of Blackwater LOM Operating Costs	1-25
Table 1-5:	Cash Costs per Ounce of Gold Produced During Periods of Mine Operation	1-25
Table 1-6:	Economic Inputs	1-27
Table 1-7:	Production Summary	1-29
Table 1-8:	Sensitivity of Economic Performance vs. Metal Prices and Exchange Rates (base case is highlighted)	1-30
Table 4-1:	Mineral Tenure Summary Table	4-5
Table 4-2:	Surface Use Permits	4-7
Table 4-3:	Details of Royalties over Blackwater Project Claims	4-9
Table 6-1:	Work History	6-2

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Table 7-1:	Drill Database Lithological Codes	7-5
Table 8-1:	Epithermal Gold Deposit Types as defined by Sillitoe and Hedenquist (2003); modified for Blackwater	8-3
Table 10-1:	Drill Hole Summary Table	10-2
Table 10-2:	Drill Contrator and Rig Type Summary Table	10-6
Table 11-1:	ALS and Eco Tech Sample Preparation Protocols	11-2
Table 13-1:	Main Metallurgical Test Programs Analyzed for PEA	13-2
Table 13-2:	Overall Gold Recovery by Oxidation State for FCL Testwork	13-9
Table 13-3:	LOM Overall Recoveries by Flowsheet Option	13-12
Table 14-1:	Density Assignments by Lithology	14-7
Table 14-2	Capping Statistics by Domain – Au	14-7
Table 14-3:	Au Grade Model Search Parameters	14-9
Table 14-4:	Ag Grade Model Search Parameters	14-9
Table 14-5:	Lerchs-Grossmann Optimized Pit Economic Assumptions	14-11
Table 14-6:	Blackwater Project Indicated Mineral Resource Summary (Effective Date July 27, 2012, Ronald G. Simpson P.Geo)	14-13
Table 14-7:	Blackwater Project Inferred Mineral Resource Summary (Effective Date July 27, 2012, Ronald G. Simpson P.Geo)	14-13
Table 16-1:	Subset of Mineral Resources Used in Proposed Production Schedule	16-1
Table 16-2:	Model Variables used in PEA Mine Plan	16-2
Table 16-3:	NSR Inputs	16-3
Table 16-4:	Input Costs for Pit Definition	16-4
Table 16-5:	Proposed Mine Production Schedule	16-9
Table 16-6:	Waste Tonnage	16-11
Table 16-7:	Waste Storage Locations	16-11
Table 16-8:	Maximum Waste Dump and Stockpile Volumes and Tonnages	16-13
Table 16-9:	Mine Production Equipment Fleet	16-20
Table 16-10:	Mine Support Equipment Fleet	16-21
Table 16-11:	Mine Supervision and Salaried Staff	16-23
Table 16-12:	Mine Hourly Personnel	16-24
Table 17-1:	Process Design Criteria – WOL Flowsheet	17-3
Table 17-2:	Reagent List	17-7
Table 20-1:	BC Provincial Permits, Licences, and Authorizations	20-9
Table 20-2:	Federal Permits, Licences, and Authorizations	20-11
Table 21-1:	Capital Cost Estimate Responsibility	21-1
Table 21-2:	Mine Capital Costs	21-3
Table 21-3:	On-Site Infrastructure Capital Costs	21-4
Table 21-4:	Process Plant Capital Costs	21-5
Table 21-5:	Major Process Plant Equipment Costs	21-5
Table 21-6:	Waste Management Capital Costs	21-6
Table 21-7:	Water Supply Capital Cost Breakdown	21-8
Table 21-8:	Indirect Costs	21-9
Table 21-9:	EPCM Costs by Area	21-9
Table 21-10:	Contingency Costs	21-10
Table 21-11:	Sustaining Capital Schedule	21-10

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Table 21-12:	Salvage Value	21-11
Table 21-13:	Initial Capital Cost Summary by Major Area	21-12
Table 21-14:	Mining Unit Costs for Different Material Types and Dump Locations	21-14
Table 21-15:	Breakdown of LOM Mining Operating Costs	21-14
Table 21-16:	Processing Costs for PEA WOL Flowsheet	21-15
Table 21-17:	Processing Labour Costs	21-15
Table 21-18:	Processing Power Costs	21-16
Table 21-19:	Grinding Media Operating Costs – WOL	21-16
Table 21-20:	Processing Reagent Use for WOL Circuit	21-16
Table 21-21:	G & A Operating Costs	21-18
Table 21-22:	Summary of Blackwater LOM Operating Costs by Cost Area	21-18
Table 21-23:	Cash Costs per Ounce of Gold Produced During Periods of Mine Operation	21-19
Table 22-1:	Economic Inputs	22-3
Table 22-2:	Royalty Requirements Covering Blackwater Deposit	22-3
Table 22-3:	Operating Costs Statistics (Unit Costs Averaged over the Mine Life)	22-3
Table 22-4:	Sustaining Capital Costs per Mine Operating Area	22-3
Table 22-5:	Production Summary	22-6
Table 22-6:	Summary Cash Flow	22-7
Table 22-7:	Sensitivities, Pre-Tax NPV at 5% Discount Rate (CDN$)	22-9
Table 22-8:	Sensitivity of Economic Performance vs. Metal Prices and Exchange Rates	22-9

Figures

Figure 1-1:	Gold Production and Cash Costs (net of royalty and silver credits)	1-25
Figure 1-2:	Selected Sensitivities:  Major Drivers of Pre-Tax NPV (CDN$)	1-30
Figure 2-1:	Project Location Plan	2-2
Figure 4-1:	Blackwater Project Location Plan	4-2
Figure 4-2:	Mineral Tenure Plan	4-3
Figure 7-1:	Regional Geology	7-2
Figure 7-2:	Geology in Vicinity of Blackwater Deposit	7-4
Figure 7-3:	Drill Hole Plan Showing Location of Referenced Cross-Sections	7-9
Figure 7-4:	Cross-Section 2800 N	7-10
Figure 7-5:	Cross-Section 5200 E	7-11
Figure 7-6:	Cross-Section 5600 E	7-12
Figure 8-1:	Schematic Section of Calc-Alkaline Volcanic Arc Setting and Associated Epithermal and Related Mineralization	8-2
Figure 8-2:	Cross-Section of Conceptual Model	8-4
Figure 9-1:	Titan-24 DC-IP Survey Lines	9-3
Figure 9-2:	Quantec 3D Results on Line 5500E	9-1
Figure 9-3:	Additional Exploration Targets Surrounding the Current Blackwater Resource	9-3
Figure 10-1:	Project Drill Hole Location Plan	10-3
Figure 10-2:	Drill Hole Location Plan, Blackwater Deposit Area	10-4
Figure 10-3:	Drill Hole Location Plan, Historic Drilling	10-5
Figure 10-4:	Loction of Metallurgical Drill Holes	10-8
Figure 14-1:	Example Section Showing Block Lithology Coding	14-2

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Figure 14-2:	Example Section Showing Oxidation Profile Coding	14-3
Figure 14-3:	Plan View, Grade Shells and Domains	14-5
Figure 16-1:	Final Pit Layout Plan	16-6
Figure 16-2:	Summary LOM NSR Cut-off Grade Profile	16-8
Figure 16-3:	Final Pit and Waste Storage Facilities Layout Plan	16-12
Figure 16-4:	Pit Design Sectors	16-16
Figure 17-1:	Process Flow Diagram for Whole Ore Leaching	17-2
Figure 18-1:	Mine Access Road and Selected Water Supply Source Location Map	18-2
Figure 18-2:	Overall Site – Layout of Open Pit and Waste Facilities	18-4
Figure 18-3:	Layout of Plant and Infrastructure	18-5
Figure 18-4:	Proposed Layout Plan, Disposal Facilities (2012)	18-11
Figure 18-5:	Project Catchment Area Boundaries	18-17
Figure 18-6:	Transmission Line Options	18-22
Figure 20-1:	Main Drainages and Watersheds in the Local Study Area	20-3
Figure 22-1:	Gold Production and US$/oz Cash Costs (net of royalty and silver credits)	22-5
Figure 22-2:	Annual Cash Flows and Cumulative Project Value (CDN$)	22-8
Figure 22-3:	Selected Sensitivities:  Major Drivers Pre-Tax NPV (CDN$)	22-9

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

1.0 SUMMARY

AMEC Americas Limited (AMEC) was requested by New Gold Inc. (New Gold) to prepare a NI 43-101 Technical Report (the Report) for the wholly-owned Blackwater project (the Project) located in central British Columbia, Canada.  The Report was prepared to support a preliminary economic analysis (PEA) on the Project, as disclosed in the New Gold news release entitled “New Gold Announces Positive Preliminary Economic Assessment at Blackwater Gold Project with Average Annual Gold Production over 500,000 Ounces at Well Below Industry Average Cash Costs” dated September 20, 2012.  All currencies are expressed in Canadian dollars (CAD$), unless otherwise stated.

1.1 Key Outcomes

The key findings of the PEA are summarized below:

· Indicated Mineral Resources at a gold equivalent cut-off grade of 0.3 g/t AuEq total 267 Mt grading 0.876 g/t Au and 4.3 g/t Ag.  In addition, there is an Inferred Mineral Resource of 120 Mt grading 0.687 g/t Au and 7.3 g/t Ag at a cut-off grade of 0.3 g/t AuEq

· Estimated operating mine life of 16.4 years:  based on conventional truck-and-shovel operation; open pit mining for first 15 years, stockpiles provide feed thereafter; strip ratio over the life-of-mine averages 2.36:1 (waste to mineralized material)

· Over the first five years of operation, the average gold feed grade will be 0.95 g/t and average annual gold production will be 569,000 oz per year

· Total LOM gold and silver production including lower grade stockpiles will be 6.2 million oz and 18.6 million oz from the Indicated category and 1.8 million oz and 13.5 million oz from the Inferred category respectively

· Process plant processes 60,000 t/a; uses conventional whole ore leach flowsheet; on average, 87% of the gold and 53% of the silver would be recovered into a gold-silver doré product

· Capital cost:  $1,814 million including a 24% contingency

· Operating cost:  $15.17/t milled; and $13.01/t after silver credits and royalty

· The PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves,

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and there is no certainty that the PEA based on these Mineral Resources will be realized.  Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability

· Net present value (base case at 5% discount rate):  $1,872 million, and US$1,655 million  (before tax, from 2015)

· After-tax net present value at 5% discount rate of US$1,123 million

· Internal rate of return:  16.4% before tax and 14.0% after tax

· Payback:  4.7 years before tax and 4.8 years after tax

· Average life of mine cash costs,  net of royalties and silver credits, of US$543/oz

· Gold will account for 93% of the revenue and silver will account for 7%.

1.2 Location, Climate, and Access

The Project lies in central British Columbia, approximately 112 km southwest of Vanderhoof and 446 km northeast of Vancouver.  The Project site is readily accessible by vehicle from the Kluskus-Ootsa Forest Service Road originating approximately 10 km west of Vanderhoof, off Highway 16.  At kilometre 147 along this road, an 18 km mine road built in 1986 and improved in 2009 provides direct access to the property and camp location.  The area is very sparsely inhabited; three ranches are found within a 20 km radius of the Project site.  Some services are available in Vanderhoof, but Prince George is the regional hub with air service from major centres.

The elevation of the Blackwater property ranges from just over 1,000 m in low-lying areas northeast of the proposed mine site to 1,800 m at the summit of Mt. Davidson on the southwest side of the property.

The climate in the project area is sub-continental.  Mining activities should be capable of being conducted year-round.

1.3 Mineral Tenure, Surface Rights, Royalties, and Agreements

The Project consists of 69 mineral claims, covering an area of 27,175 ha.  A total of 66 of the claims are recorded in the name of New Gold.  The remaining three claims, 637203, 637205, and 637206, termed the JR Claims, are recorded in the name of third parties.  New Gold was granted an option to acquire a 100% interest in the JR Claims, subject to a royalty.

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The currently-defined Blackwater deposit spans one of the three Davidson claims (509273, but not 509274 and 509275), the Dave claim (515809) and the Jarrit claim (515810).  Future development of the mineral claims covering the Blackwater deposit will require conversion of those claims to mining leases.

Five royalties are payable on production from the New Gold property holdings; however, only three royalties affect the Blackwater deposit.  The royalties are a 2% net smelter return (NSR) on mineral claims 509273, 509274, and 509275; a 1.5% NSR on mineral claim 515809, a 1% NSR on mineral claim 515810, a 3% NSR on mineral claims 637203, 637205, and 637206, and a 2% NSR on mineral claim 835014.  New Gold may purchase the 2% NSR on mineral claims 509273, 509274, and 509275 for $2 M, 2% NSR of the 3% NSR on mineral claims 637203, 637205, and 637206 for $1 M (at any time after it exercises the option to acquire a 100% interest in those claims), and 1% NSR of the 2% NSR on mineral claim 835014 for $1 M.  The economic analysis assumes that the NSR on mineral claims 509273, 509274, and 509275 will be bought out, and the only remaining payable royalties for the Blackwater deposit will be the 1.5% NSR on mineral claim 515809, and 1% NSR on mineral claim 515810.

The claims are located on Crown land.  Limited searches that have been conducted did not reveal any dispositions of surface rights by the Crown under the Land Act within the Project area, although surface use permits, licences and authorizations such as traplines and licences-to-cut are active.  There is a reasonable expectation that with appropriate supporting studies as required by permitting authorities that land access and provision of land for infrastructure development for any proposed mining activity will be achievable.  There is sufficient suitable land area available within the mineral claims for any future tailings disposal, mine waste disposal, and installations such as a processing plant, and related mine infrastructure.

1.4 Environment, Permitting, and Socio-Economics

Baseline studies have been initiated and are ongoing.  Studies to date have included species at risk, a terrestrial ecosystem mapping (TEM) survey, and the collection of plant samples for baseline metal concentrations, evaluation of water quality samples, and wildlife surveys within and adjacent to the project footprint.

Water quality monitoring is ongoing to establish background conditions.  To date, dissolved aluminum and total cadmium arising from natural sources have exceeded guidelines at a number of sites.

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Culturally-modified trees were identified in one location near the Project area.  No other archaeological sites, historic heritage sites, or evidence of any previous archaeological investigations were identified within the project area.

The Project is subject to review under the BC Environmental Assessment Act and the Canadian Environmental Assessment Act, 2012.  A Project Description to initiate the EA process, has been written in parallel with completion of the PEA, and is being filed in late 2012.  Following EA approval, additional permits, licences, authorizations, and certificates will be required to proceed to construction of the project.  New Gold has identified and commenced planning for obtaining all such approvals in parallel with the EA approval process.

Draft Terms of Reference for the provincial Environmental Assessment Certificate Application (EAC) and draft guidelines for the federal Environmental Impact Statement (EIS), are expected to be filed in early 2013 for approval by the BC Environmental Assessment Office (BCEAO) and Canadian Environmental Assessment Agency (CEA Agency).  An EAC Application/EIS approval will then be filed, and reviewed with the applicable agencies, First Nations and stakeholders.  Following completion of this review, New Gold anticipates receiving required Environmental Assessment (EA) approval in mid to late 2014.  Following EA approvals, additional permits, licenses, authorizations and certificates, will be required to proceed to construction of the project.  New Gold has identified and commenced planning to obtain all such approvals in parallel with the EA approval process.  These latter approvals are expected by early 2015.

The Project is subject to federal and provincial permitting requirements, which include public and First Nations consultations.  The BCEAA and CEAA, 2012 have specific, regulated timelines, and in addition, provincial and federal permits have specific timelines and requirements that must be met.  New Gold has outlined a permitting schedule based on these timelines; however changes to the schedule may occur as a result of the consultation processes and any subsequent amendments that may be then required to support furtherance of permit application processes.

There has been no evidence of any historic mining activity identified on any of the claims to date.  A Phase 1 Environmental Site Assessment was conducted for the proposed mine site, and no records of site activities were identified prior to 1973.  It is confirmed that the potential for environmental liabilities is limited to exploration activities conducted since 1973.  These potential liabilities are within the reclamation bond amount.

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Existing environmental liabilities are related to the exploration activities that New Gold has undertaken to date or to exploration activities conducted by third parties prior to New Gold’s acquisition of the Blackwater mineral claims.  A reclamation plan for the exploration activities was prepared to accompany the Notice of Work and Reclamation application to the BC Ministry of Energy and Mines, and addresses drilling, geophysical surveys, access to site, and the camp and camp services, and proposes a series of measures that include removing buildings, machinery, debris, or contaminated soils; re-establishing natural landscapes; and deactivating roads and trails.  The current established security deposit for project reclamation is approximately $1 million.

A key environmental consideration for the project is the presence of fish-bearing habitat.  Fisheries values are higher in the lower reaches of Davidson Creek, where Kokanee salmon spawning habitat has been identified. The project area avoids these lower reaches.

New Gold has made contact with First Nations groups who have interests in the Project area, and the company is engaged in ongoing consultation with Provincial and Federal ministries and agencies about research design, permits, and environmental assessment questions and issues as they arise.

Mine closure costs for the PEA mine plan and infrastructure as envisaged in this Report are estimated at $101 million.  This cost includes reclamation and closures of the tailings storage facility (TSF) and waste dumps, decontamination of facilities as required, removal of all constructed surface facilities, removal of the power and pipelines, closure and reclamation of access roads, and post-closure monitoring and maintenance to ensure that end land use objectives are met.

New Gold has signed a Road Maintenance Agreement with Canfor Forest Products Ltd. for the use of 150 km of the Kluskus Forest Service Road to access the project, but it is only responsible for payment of a portion of the maintenance costs.  No existing environmental liabilities along the road are the responsibility of New Gold.

1.5 Geology and Mineralization

The Blackwater deposit is considered to be an example of an epithermal-style deposit.

Due to lack of outcrop, all geological interpretation has been based on drill information.

Mineralization at Blackwater is hosted within felsic to intermediate composition volcanics that have undergone extensive silicification and hydrofracturing in association with pervasive stockwork veined and disseminated sulphide mineralization.

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Gold-silver mineralization is associated with a variable assemblage of pyrite-sphalerite-marcasite-pyrrhotite ± chalcopyrite ± galena ± arsenopyrite (± stibnite ± tetrahedrite ± bismuthite).  Hydrothermal alteration (and possibly contact metamorphism) has produced several superimposed alteration assemblages, including pervasive silica-sericite-clay (illite) ± biotite alteration and veinlet/fracture-controlled silica-sericite-chlorite-clay ± iron carbonate ± tourmaline.  An early (?) biotite-silica-albite ± chlorite/actinolite hornfelsing event may have been significant, although mineralization in these rocks appears to be lower than in units without evident hornfelsing.  Visible native gold has been noted in some drill cores.

Disseminated Au-Ag mineralization is defined by an east-west trending tabular-conical shaped zone with a lateral extent of up to 1,250 m east-west x 1,000 m north-south.  Mineralization remains open at depth in the southwestern portion of the deposit as well as to the north and northwest.  The centre of the deposit has an average thickness of 350 m with a vertical extension of up to 600 m where open.  The overall volume plunges shallowly to the north and northwest with inferred steep, north-plunging higher-grade mineralized shoots measuring tens of metres thick, likely influenced by near-vertical structural intersections.

In the opinion of the QPs, knowledge of the deposit setting, lithologies, structural and alteration controls on mineralization, and mineralization style are sufficient to support Mineral Resource estimation.

1.6 Exploration

Work by parties other than New Gold has included claim acquisition, geochemical and geophysical surveys, line cutting, trenching, prospecting and reconnaissance geological mapping and reverse circulation (RC) and core drilling.  Work was completed by Granges Inc (Granges), Silver Quest Resources Ltd. (Silver Quest), Geo Minerals Limited (Geo), and Richfield Ventures Corp. (Richfield).

Since acquisition of Silver Quest, Geo and Richfield in 2011, New Gold has completed additional core drilling, and undertaken preliminary mining studies, culminating in the PEA that is the subject of this Report.

In the opinion of the QPs, the exploration programs completed to date are appropriate to the style of the deposits and prospects within the Project.  The exploration and research work supports the genetic and affinity interpretations for the deposit.

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1.7 Exploration Potential

The Blackwater project area offers good potential for the discovery of additional mineralization that may support mineral resource estimation.  Work to develop this potential is ongoing and involves a combination of detailed interpretation of deposit geology from the growing body of exploration drill hole information in conjunction with expanded geologic mapping, geochemical sampling, airborne and ground-based geophysical methods, and exploration drilling, both beyond the limits of the Blackwater deposit as they are currently known, and within the greater project area.

Within the Blackwater deposit itself, potential exists to expand the known mineralization along the western and northwestern margins.

Elsewhere within the project area, surface geochemical sampling and exploration drilling have recently identified an area of favorable alteration and mineralization approximately 2 to 2.5 km northwest of the Blackwater deposit. Geochemical sampling has identified an area of anomalous gold-in-till values. Additionally, drilling in this area has intercepted rocks with similar lithologies, styles of alteration and mineralization to those observed in the Blackwater deposit.  These areas warrant further exploration.

1.8 Drilling

A total of 449 drill holes (168,276 m) have been drilled in the Project area between 2009 and May 2012.  Of this total, 134 were completed by Richfield, and 315 by New Gold.  An additional 34 RC holes were completed by Granges; however the precise locations of these drill holes are uncertain.  Drilling is ongoing on the Project.

Exploration drilling carried out since 2009 has been predominantly HQ diameter (63.5 mm) diamond drill core except where a reduction to NQ diameter (47.6 mm) was required to attain target depths.  The last three metallurgical holes (BWMET05 – BWMET07) were PQ diameter (85 mm) core.  Core recovery for the 2009, 2010, and 2011 drilling programs averaged 91%, and the median core recovery was 97%.  Poor core recovery often occurs in zones of faulting and fracturing.  The average RQD value for this period was 44%.

Planned drill hole collar locations were measured in the field using hand-held global positioning system (GPS) instruments.  Locations were subsequently confirmed by Trimble differential GPS.  Elevations for the drill collars were determined by draping collar coordinates over the topography measured by a 2011 LiDAR survey.  Down-hole surveys are performed using Reflex survey equipment, and dip angle and azimuth are recorded.  A +18.8° magnetic declination correction factor is applied to the magnetic azimuth record.

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Logging includes geotechnical, magnetic susceptibility, and specific gravity measurements taken at regular intervals.  Lithology is logged and the core prepared for systematic sampling at regular 1 m intervals.  The lithological nomenclature at the Project has undergone revision on two occasions since New Gold took control of the Project in June of 2011, and currently six principal rock lithology types are defined.

Magnetic susceptibility and conductivity data were measured at 10 cm increments along the core with a hand-held conductivity and magnetic susceptibility meter.

In the opinion of the QPs, the quantity and quality of the lithological, geotechnical, collar and down-hole survey data collected in the exploration and infill drill programs completed by Richfield and New Gold from 2009 to 2012 are sufficient to support Mineral Resource estimation.

1.9 Sample Analysis and Security

Drill core is transported in core boxes from the drill site to camp by four-wheel drive vehicle and logged in a specially built core handling facility, where sampling for analysis is done systematically on 1 m sample composites.  Intervals are measured and marked on the core boxes, where sample tags are stapled at the beginning of each interval.  A diamond saw is used to split the core in half lengthwise.  One half of the core is placed in a standard heavy poly sample bag identified with a re-printed sample tag in a zip-lock plastic bag, and the poly bag is closed with a zip tie.  The remaining half core is kept for reference in the core box and stored at the project camp site.

Certified reference standards (CRMs), blanks, and duplicates are inserted into the sample stream, and four sample bags are then placed into a larger rice bag labelled with the sample identification (ID) numbers and sealed with a numbered banker’s security tag.

Eco Tech, the primary analytical laboratory used by Richfield, had ISO 9001:2008 certification.  ALS Minerals (ALS), which has been used as the primary analytical laboratory since July 2011, when the company acquired Eco Tech, is ISO 9001:2008 accredited, and has ISO/IEC 17025:2005 accreditations for specified analytical techniques.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Standard preparation procedures are used to crush, split, and pulverize the core samples before analysis.  Eco Tech crushed samples to 70% passing 2 mm, riffle split 250 g, and then pulverized the split to 95% passing 106 µm.  ALS crushed samples to 90% passing 2 mm, took a 1,000 g split, and pulverized the split to 95% passing 106 µm.  In late 2011 and into 2012, approximately 7,086 samples were prepared at ALS in Elko, NV, USA.  This amount represents approximately 5% of the total sample data set used for the mineral resource estimate.  The Elko samples were pulverized to 95% passing 75 µm.

Assay procedures include a multi-element package (28 elements at Eco Tech, 33 elements at ALS) with aqua-regia (AR) digest (nitric and hydrochloric acids) and induction-coupled plasmaspectrometry atomic emission spectrometry (ICP AES) finish for multi-element analysis.  Both laboratories used fire assay with atomic absorption (AA) for gold, although those performed by Eco Tech were based on 30 gram pulp splits whereas those performed by ALS were based on 50 gram pulp splits.  In earlier samples, Eco Tech assayed silver by aqua regia/AA, but later switched to include silver in the overall ICP analysis.  ALS uses ICP for silver analysis.

Because of the differences in assay method and some documented discrepancies in sample identification, caution is needed in interpreting the accuracy of silver grades relative to gold grades reported in the current mineral resource estimate.

At both laboratories, overlimit ICP analyses of Ag, Cu, Pb, and Zn are conducted on analyses that returned grades of >30 g/t Ag and >1% Cu, Zn, and Pb.  For Au, overlimit fire assay (FA) analysis is conducted on analyses that return >10 g/t.

QA/QC protocols for the Richfield work included the insertion of certified reference material (CRM), a blank, and a quarter-hole duplicate into every set of 30 samples.  The CRMs were produced by CDN Resource Laboratories Ltd., an independent laboratory in Langley, BC.  The blanks inserted into the sample stream were originally derived from an outcrop of barren granodiorite exposed about 10 km north of the mineralized zone.  Eco Tech analyzed 13 samples of this material in 2009 before using it for sample blanks.  Limestone garden stone was later used for the blanks.

In the New Gold QA/QC program, one CRM standard is inserted every 20 samples, and one blank is inserted every 30 samples.  As of October 2011, quarter-core field duplicates are no longer inserted.  The reference standards acquired from CDN Resources were used until the end of 2011 and were phased out of the program by February 2012.  Since then, CRMs have been sourced from from WCM Minerals, a division of WCM Sales Ltd. in Burnaby, BC, and Geostats Pty, in WA, Australia.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

The drill hole database is supported by some 40,000 quality assurance/quality control (QA/QC) check assays.

The current drill hole and assay database for the project is administered from the New Gold Vancouver office using the same Maxwell GeoServices sortware products, LogChief™ and DataShed™, which were used during the Richfield drilling programs.

1.10 Data Verification

Data verification has consisted of site visits to inspect drilling, sampling, and quality assurance/quality control procedures.  The geology and mineralization encountered in the drill holes completed to date were also reviewed.

QC/QC data in the form of blanks, CRMs, check assays and duplicate samples have been reviewed.  QA/QC with respect to the results received to date for the 2009, 2010, 2011 and 2012 exploration programs is acceptable, and protocols have been well documented.  New Gold is currently investigating a possible decrease in analytical precision noted in analytical QA/QC samples analyzed by ALS since July 2011.

In November 2011 and January 2012, AMEC completed data verification of the drill hole database used to support New Gold’s Blackwater mineral resource estimate.  The database verification was performed to confirm that drill hole data used to support the resource estimate are accurately transcribed.  Transcription accuracy was assessed through comparison of the drill hole database records to original collar location, down-hole survey, and assay documentation where available.  As a result of the database verification, AMEC concluded the drill hole database is reasonably free of transcription errors.  Although some errors and deficiencies were observed, AMEC concluded these would not have a material impact on the resource estimate.  The AMEC database verification work and recommendations to address observed errors or deficiencies are documented in an internal memorandum to New Gold.

The process of data verification performed by the QPs indicates that the data collected from the Project adequately reflect deposit dimensions, true widths of mineralization, and the style of the deposits, and adequately support the geological interpretations, and the analytical and database quality.  Drill data are typically verified prior to Mineral Resource estimation by comparing data in the Project database to data in original sources.  For most of the data, the original sources are electronic data files; therefore, the majority of the comparisons were performed using software tools.  No significant errors were found with the database that would preclude use in Mineral Resource estimation.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

1.11 Metallurgical Testwork

A metallurgical testwork program for the study was carried out on samples that were composited to represent process plant feed in the mine development plan.  These samples were obtained from two primary sources:  a dedicated four-hole metallurgical drilling, and composites from exploration drilling of up to 89 holes.

The 2011-2012 test work program included the following stages:

· Comminution parameters on samples from different lithologies, including Bond work indices, and JK SAG mill parameters

· Bottle-roll cyanide leaching tests of ground oxide, transition and sulphide materials for a potential milling and whole ore leach flowsheet, and at a range of crush sizes for a potential heap leach flowsheet

· Column leach work on crushed transition and sulphide mineralized materials

· Batch flotation tests, primarily on sulphide material of different lithologies, including cyanide concentrate leaching on a sub-sample of tests

· Indicative gravity separation tests, including the interaction of gravity with flotation and leaching test circuits

· Settling and filtration tests on ground mineralized material and concentrate slurries.

Primary areas of investigation for the whole ore leach program included primary grind size and leach retention time.  Primary areas of investigation for the flotation and concentrate leach program included primary grind size, pulp pH, reagents, flotation retention times on the rougher/scavenger flotation circuits and regrinding, leach retention time and cyanide concentration on leaching of rougher/scavenger flotation concentrates.

The test work results were used to determine the relationship between oxidation state and recovery and to project the following life-of-mine average metallurgical recoveries for whole ore leach:

· Gold recovery – 87%

· Silver recovery – 53%

In the opinion of the QPs, the metallurgical test work completed on the Project has been appropriate to establish a process route that is applicable to the mineralisation types.  Tests were performed on samples that were representative of the mineralisation for the purposes of establishing conceptual process flowsheets.  

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

The process routes proposed all use conventional technology.  Recovery factors from the tests are appropriate to the mineralisation types, based on the available metallurgical testwork data.  On a daily basis, recovery will vary depending on feed grade and mix of mineralization types but it is expected that typically the recovery will follow the average values assumed in the study.

Overall results indicate that the gold in Blackwater mineralization is free-milling (non-refractory).  A portion of the gold, between 17% and 32%, liberates during grinding and is amenable to gravity concentration.  In addition, a preliminary diagnostic leach test indicated that 90% to 95% of the overall gold is cyanide-soluble, with only 1% to 5% of the gold occluded in sulphides.  Silver, despite higher grades, exhibited poor amenability to low mass pull gravity concentration, and diagnostic leach tests showed moderate dissolution, in the range of 55% to 62%.  This level of dissolution is in line with observations of silver minerals such as friebergite, which has been reported to be refractory.

1.12 Mineral Resource Estimate

A total of 449 holes totalling 168,276 m and 144,301 sample assays were used to support the grade estimations.  Data from the remaining holes were used in the density and lithologic modelling.

Lithological wireframe models of the main fragmental, andesite, and sedimentary units were generated by company geologists using cross-section interpretations.  A bedrock surface was modelled by creating profiles based on drill hole intercepts and generating a digital elevation model.  Surfaces representing the base of the weathered oxide and oxide-sulphide transition zones were also generated.  The deposit was also divided into two structural domains based on large-scale gold trends.  Gold grade-shell domains were then generated within each structural domain by indicator kriging using a threshold value of 0.2 g/t Au.  A higher-grade Ag domain (GS10) lies along the north side of the deposit close to surface.  An indicator-kriged grade shell using a threshold value of 10 g/t Ag was generated as a domain.

A total of 18,792 density measurements made on core sampled between 2009 and June 8, 2011 have been used for density modelling.  Separate interpolations were carried out within the oxide, transition, and sulphide zones using hard boundaries.  Soft boundaries were used between the andesite and fragmental lithologic domains.  An isotropic search ellipsoid was used with a maximum search distance of 150 m.  A minimum of two and maximum of nine samples were required to estimate a block.  A maximum of three samples was permitted from a single drill hole.  Blocks that were not estimated were assigned the mean value for the corresponding lithology and material type.

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Grade caps were used for gold and silver assay values.  After reviewing the probability distribution, it was decided to cap Au grades at a level of 40 g/t.  The capping affects a total of 62 samples and removes an expected 4.3% of the metal.  A cap grade of 150 g/t was selected for Ag within the structural domains but outside of the 10 g/t Ag grade shell domain.  This affects 74 samples and removes and expected 2.9% of the metal.  Silver samples within the 10 g/t Ag grade shell domain were not capped.

Down-hole composites for Au were created within the domains using the “best fit” method.  This procedure produces samples of variable length but of equal length within a contiguous drill hole zone, ensuring the composite length is as close as possible to the nominated composite length.  In this case, the nominated length was set at 10 m.  Fixed-length down-hole sample assay composites of Ag were created using a 10 m interval within the broader structural domains.  Samples were separated into two populations, those within a 10 g/t Ag grade shell domain and those outside of the domain.

To determine kriging parameters, search parameters, and anisotropy, directional pairwise relative semi-variograms for Au and Ag were modelled using composites falling within the corresponding domain constraints.  In the case of Au, composites were restricted to the two indicator grade shells.  For Ag, the composites were combined into a single population, as the number of sample pairs within the 10 g/t Ag grade shell was inadequate to model variograms.

A block model was created in Gemcom-Surpac Vision© software using a block size with dimensions of 10 m x 10 m x 10 m.  The model blocks were first coded by the partial percent within the zone domains and below topography.  Lithologic codes, material types, and density values were then assigned.  Indicator kriging was carried out using Leapfrog3d© software to generate grade shell solid models within each of the two structural domains at a threshold of 0.2 g/t Au.  Gold and Ag grades within the corresponding zone domains were estimated in three passes using ordinary kriging and the inverse distance squared weighting method (ID2).  A single pass nearest-neighbour estimate was also carried out for use in model validation.

Review of the models and interpolation indicated no significant biases or errors.

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Consideration was given to classifying blocks estimated in the first estimation pass as Measured.  However, the structural and geologic modelling is still at an early stage, and some assay/data quality issues are still under investigation.  Consequently, it was decided to delay classifying any blocks as Measured for the current estimate.  Blocks were initially classified as Indicated if the block centroid was within 38.5 m of the closest hole, the second closest hole was within 50 m, and the average distance from the two closest holes was also less than 38.5 m.  This corresponds to a drill spacing of 50 m x 50 m with a 10% contingency buffer.  The Indicated classification was then smoothed to eliminate small areas of Inferred surrounded by Indicated blocks.  Finally, Indicated blocks adjacent to the outer grade-shell boundaries were downgraded to Inferred to reflect the uncertainty of grade estimates along the hard boundaries.

To assess reasonable prospects for eventual economic extraction a Lerchs-Grossmann (LG) optimized pit was prepared using general economic and technical assumptions appropriate to the deposit style and location.  A gold equivalent (AuEq) grade was calculated using assumed metal prices of US$1,300/oz and 83.6% metallurgical recovery for gold and US$24/oz and 44.9% metallurgical recovery for silver.  A 0.23 g/t gold-equivalent cut-off would return an estimated $8.06/t for Blackwater open pit mineralization.  This covers the assumed processing and general and administrative costs, and represents a marginal cut-off grade.  The metallurgical recoveries used in the LG assumptions were based on preliminary data and differ from the updated recovery figures used in the economic analysis.

1.13 Mineral Resource Statement

Mineral resources have been classified using the definitions set out in CIM (2010).  Mineral resources have an effective date of July 27, 2012.  The QP is Ronald G. Simpson P.Geo.  Mineral resources are summarized in Table 1-1 and Table 1-2, which presents the basecase estimate at an AuEq cut-off grade of 0.3 g/t AuEq, and sensitivity estimates around this basecase to show the sensitivity of the mineral resource estimate to changes in AuEq grades.

The distribution of resource tonnes by material type for the base case is 3% within the oxide zone, 9% within the transition zone, and 88% in the sulphide zone.

These data differ slightly from those reported in the New Gold July 18, 2012 News Release because the LG resource pit shell and gold equivalent cut-off were based on an assumed silver price of US$23/oz, which has been increased to US$24/oz for this resource statement.  This silver price change marginally changes the gold equivalent formula.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Table 1-1: Blackwater Project Indicated Mineral Resource Summary (Effective Date July 27, 2012, Ronald G. Simpson P.Geo)

Cut-off Grade	Tonnes 000's	Grade				Contained Metal
AuEq g/t		Au g/t	Ag g/t	AuEq g/t		Au M oz	Ag M oz	AuEq M oz
0.20	285,241	0.836	4.2	0.878		7.67	38.52	8.05
0.25	280,379	0.847	4.2	0.889		7.64	37.86	8.01
0.30	267,145	0.876	4.3	0.919		7.52	36.93	7.89
0.35	249,371	0.917	4.5	0.962		7.35	36.08	7.71
0.40	230,572	0.963	4.6	1.009		7.14	34.10	7.48
0.45	210,646	1.017	4.7	1.064		6.89	31.83	7.21
0.50	191,572	1.074	4.9	1.123		6.61	30.18	6.92

Table 1-2: Blackwater Project Inferred Mineral Resource Summary (Effective Date July 27, 2012, Ronald G. Simpson P.Geo)

Cut-off Grade	Tonnes 000's	Grade				Contained Metal
AuEq g/t		Au g/t	Ag g/t	AuEq g/t		Au M oz	Ag M oz	AuEq M oz
0.20	133,219	0.642	7.0	0.711		2.75	29.98	3.05
0.25	128,572	0.659	7.0	0.728		2.72	28.94	3.01
0.30	120,478	0.687	7.3	0.759		2.66	28.28	2.94
0.35	110,383	0.724	7.5	0.798		2.57	26.62	2.83
0.40	98,937	0.771	7.8	0.848		2.45	24.81	2.70
0.45	88,061	0.822	7.9	0.900		2.33	22.37	2.55
0.50	77,525	0.882	7.7	0.958		2.20	19.19	2.39

Notes to accompany Mineral Resource tables:

1.	Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability
2.	Mineral Resources are amenable to open pit mining methods and have been constrained using a Lerches-Grossmann optimized pit.
3.	Assumptions include US$1,300/oz Au, US$24/oz Ag, US$1.00:CAD$1.00, 83.6% Au recovery, 44.9% Ag recovery, $1.52/tonne mining cost, $1.90/tonne waste mining cost, $10.52/tonne process and G&A cost.  No allowances have been made for mining losses and dilution. The average pit slope angle is assumed to be 40°.
4.	The base case gold equivalent (AuEq) cut-off (bolded) is greater than the conceptual marginal cut-off of 0.23 g/t.
5.	AuEq = US$ 24/oz Ag x 44.9% / US$ 1,300/oz x 83.6%
6.	Gold analyses are performed by fire assay/AA finish methods and silver analyses are performed by Induction Coupled Plasmaspectrometry (ICP). Silver ICP analyses are not known with the same precision and do not have the same quality control support as gold fire assay analyses
7.	Rounding as required by reporting guidelines has been used, and totals may not sum.

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Areas of uncertainty that may materially impact the Mineral Resource estimates include:

· Commodity price assumptions

· Assumptions that required permits will be forthcoming

· Assumptions used in the LG conceptual pit shell, including pit slope angles, metal recovery assumptions, and mining, general and administrative and process cost assumptions

Relogging of older drill holes may also result in local changes to the lithological interpretations on which the current block model and therefore the resource estimate, is based.

1.14 Drilling Completed Since Database Closeout Date

Drilling on the Project is ongoing.  From the PEA resource cut-off up to September 14, 2012, 377 additional holes and 100,363 m of drilling has been completed.  Complete or partial analytical results for an additional 76 holes (27,968 m) were received between the closeout date and August 28, 2012.  Of these, 69 holes (24,892 m) were infill and seven (3,076 m) were step-out or boundary definition holes.  A review of the drill data available as at August 28, 2012 was performed to compare the results with the block model.  Although incorporation of these drill holes in an updated block model will result in local differences in interpretation of the geological and grade shells, visual inspection of the drill holes against the block model indicated the results were generally consistent with the current interpretations as used in the model.

A review of drill log information from post database cut-off holes for which assays had not been received was also carried out.  No obvious geological features were identified that would significantly change the interpretations used to construct the block model that supports the mineral resource estimate.

1.15 Proposed Mine Plan

1.15.1 Pit Design and Scheduling

The mining section is preliminary in nature and partly based on Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the PEA based on these Mineral Resources will be realized.  Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Conventional open pit mining will be employed to mine the Blackwater deposit.  The mine plan includes both Indicated and Inferred Mineral Resources to make up the mill feed tonnage.  Material not processed is waste.  The life-of-mine (LOM) mill feed production will be 255.9 million tonnes (Mt) from the Indicated category and 100.0 Mt from the Inferred category. Mill feed will be delivered to a primary crusher that feeds a whole ore leach (WOL) gold–silver recovery plant.  The milling rate is 60,000 metric tonnes per day (t/d), or 21.9 Mt per year (Mt/a).  The mine life is 16.4 years after commercial production begins.  The total tonnage mined from the open pit is 1,195 Mt at an initial daily mining rate of 169,000 t/d and increasing to 273,000 t/d during Years 3 through 10 before declining after the peak waste mining period.  A total of 7.2 million ounces of gold and 35.3 million ounces of silver are estimated to be contained in the Indicated category of mill feed tonnage during the life of the mine.  An additional 2.1 million ounces of gold and 25.5 million ounces of silver are estimated to be contained in the Inferred category of mill feed.

Some of the mill feed, referred to a low-grade mineralized material, will be stockpiled in the early years for re-handle into the mill at the end of the mine life.  Another sub-set of the mill feed is within a high silver–low gold area in the northern part of the deposit, referred to as high-silver mill feed.  This material is also planned to be stockpiled and re-handled to the mill.

An NSR value is calculated for each block in the model with an assigned gold grade.  The NSR value is used to express the value of a block because both gold and silver contribute to the value and their recoveries vary by the mill feed type (based on oxidation state), and because different royalties apply to the various claim blocks that host the deposit.  The NSR value therefore incorporates the gold and silver grades, the unique metal recoveries in the mill, metal prices, the various royalties, and post-property doré costs.  The high-silver zone in the northern part of the deposit has been identified as a unique mill feed because of the high silver grades (average 45.5 g/t in mill feed) and low gold grades (average 0.067 g/t).  The recovery of gold at these low grades has not been tested, thus for the PEA mine scheduling purposes, the gold recovery is set to zero for gold values less than 0.07 g/t.  However, in the economic analysis, a 20% gold recovery has been assigned to the material.

The open pit was optimized at a $9.20/t NSR cut-off value as determined from costing for milling, plant services, tailing services and general and administrative (G&A) charges and at US$1,200/oz gold, US$20/oz silver and 1.00 US$/CAD$ exchange rate.  The mine production schedule incorporates an elevated cut-off grade strategy during the first seven years to raise the mill feed grade.  Material below that cut-off grade and above a $10.00/t NSR cut-off is stockpiled for processing at the end of the mine life.  The $10.00/t NSR cut-off is selected (versus the $9.20/t NSR used for pit optimization) for stockpiling in order to cover costs for material re-handle from the stockpile and to account for a projected lower recovery of metal from stockpiled material.

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A suite of floating cone runs based on percent of NSR were generated to evaluate the development of the pit from a cone based on 40% of NSR value to 100% of NSR value.  The net value contained within the cone geometry continues to increase up to 90% of NSR, where it flattens off with little increase in value per tonne as the cone shell increases in size.  This is in part because of the increase in waste tonnage per tonne of mill feed.  The 90% of NSR cone geometry was selected as the guide for the final pit design geometry.  The final pit is sub-divided into four mining phases based on the progression of the floating cone geometries as the percent of NSR value is increased.  During the phase and final pit design process, haul roads are added and pit walls are smoothed compared to the floating cone geometry.  The final pit design includes 98.4% of the cone mill feed tonnage and 3.3% additional waste tonnage.

Phase 1 is located in the central pit area and mines the shallower high-grade mill feed with the pit bottom at the 1430 elevation.  Phase 2 expands Phase 1 in all directions (except to the northeast), and the pit bottom is lowered to the 1300 elevation.  Phase 3 expands the pit to the north, picking up the north high-silver zone and a deep high-grade zone with the pit bottom of Phase 3 at the 1170 elevation.  Phase 3 is the final pit limit in the north-central part of the pit.  Phase 4 extends the pit out to the final pit limits on the east, south, and west walls, and the pit bottom is at the 1130 elevation.  The proposed open pit will reach a maximum depth of 400 m.

The planned mine schedule starts with 30 million tonnes (30,000 kt) mined during a preproduction period, of which 5,783 kt of mill feed is stockpiled for processing in Year 1 and 1,202 kt of low-grade mill feed (NSR value greater than $10.00 and less than $15.00/t) is stockpiled.  The 23,015 kt of waste is partly used for TSF construction.  During Year 1, 19,710 kt of mill feed is processed (90% of mill capacity to allow for a ramp-up period), including the mill feed stockpiled during the preproduction period.  The high silver–low gold mill feed material located in the north side of the Blackwater deposit and mined as part of Phase 3 is stockpiled during Years 3 through 7.  This tonnage amounts to 9,975 kt and is processed during Years 8 through 14 at a rate of 1,500 kt/a.  Stockpiling of this material reduces the impact of the 0.069 g/t average gold head grade of this material on the overall gold mill head grade during Years 3 through 7.

1.15.2 Geotechnical Considerations

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Knight Piésold provided design criteria for overall and inter-ramp slopes for the pit walls.  The inter-ramp slopes are 27° in the overburden and range from 38.5° to 46°, in rock, depending on the pit wall orientation.  The overall slope angles (including flattening for the overburden) range from 34° to 43°, depending on the pit design sector.  Preliminary pit slope recommendations are based on a review of the currently available geomechanical information, stability analyses results, assumptions on mining equipment, and experience at similar open pit operations in the Project region.

1.15.3 Equipment

Mine equipment requirements were calculated based on the annual mine production schedule, the mine work schedule, and equipment shift production estimates.  The size and type of mining equipment is consistent with the size of the Project: peak run-of-mine material movements of 97 Mt/a.  Equipment represents a conventional truck-and-shovel fleet.

Mining operations will be carried out with an initial equipment fleet (during Year 1) comprising two 311 millimetre (mm) electric blast hole drills, two 56 m3 electric cable shovels, one 18 m3 front-end loader and sixteen 290 t trucks.  These will be supplemented with back-up graders, and track and rubber-tired dozers.  A 10 m bench height has been selected for mining both mineralized material and waste.  The total tonnage increases to 97 Mt/a in Years 3 through 10 and the mining fleet increases with the addition of two blast hole drills, one electric cable shovel, and a maximum addition of 19 haul trucks.

In the opinion of the QPs, equipment and personnel requirements have been considered in relation to the mine plan and are appropriate for the proposed mine production rate, sequencing, and mine life.

1.16 Process Design

The 60,000 t/d process plant will utilize conventional crushing, grinding, leaching and CIP to produce a gold-silver doré.  The overall design philosophy was to select large equipment and to maintain a simple and conventional flowsheet.

Run-of-mine (ROM) mineralized material will be crushed to 80% passing 120 mm and then ground to 80% passing 150 µm.  A gravity circuit consisting of centrifugal concentrators will treat a portion of the primary cyclone underflow to recover coarse metallic gold.  The gravity concentrate will be directed to an intensive cyanidation reactor for extraction of gold and silver.

Ground mineralized material will be directed to a leach feed thickener, then to a leaching and carbon-in-pulp (CIP) extraction circuit.  Extracted gold and silver will be released from carbon in the stripping columns and recovered by electrowinning before being smelted into a gold-silver doré product.  Stripped carbon will be treated in a regeneration kiln.

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Key process equipment will consist of:

· A primary crushing plant with a 1520 x 2260 mm (60" x 89") gyratory crusher

· A SAG/ball mill/crusher grinding circuit:

- two 11.0 m diameter x 7.3 m long (36' x 24') 16-MW SAG mills

- two 7.9 m diameter x 12.8 m long (26' x 42') 16-MW ball mills

- two 1,000 kW pebble crushers

· Whole ore leaching and carbon-in-pulp circuit:

- 24 leach tanks of 18 m diameter

- six CIP tanks of 18 m diameter

Leach CIP residue will report to cyanide destruction using the SO2-Air process and on to the TSF as a single stream.

1.17 Waste Management

Five types of waste are currently classified at Blackwater:  overburden, two types of non-acid generating rock (NAG4 and NAG3), and two types of potentially acid generating rock (PAG2 and PAG1).  Some of the material types will be used for the construction of the tailings storage facility (TSF) embankments during the mine life, and the rest will be stored in specified locations near the planned pit.  Waste dumps and stockpiles have been designed to hold all waste materials and low-grade mill feed tonnages not delivered to the primary crusher or the TSF facility.

Geotechnical design of the TSF is supported by preliminary geotechnical site investigations.  It has been designed to contain 509 Mm3 of material and will require 78 Mm3 of construction material, about 97% of which will be waste rock and overburden from the open pit.  Use of open pit materials is fully integrated into the mine plan and will minimize conventional waste rock storage facilities.  A total of 101 Mm3 of material will be stored in engineered waste rock facilities outside of the TSF.

The TSF embankments will be an engineered, water retaining, zoned earthfill/rockfill dam with a compacted low-permeability till core and appropriate filter/transition zones.  A total of three embankments will be utilized across the two sites.  The TSF dams will be expanded using the centreline construction method.

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A preliminary site water balance has indicated a net water input requirement to provide sufficient process water and to maintain a water cover over PAG waste rock and tailings in the TSF facility during some years of operation.  It is proposed that additional water to support operations will come from Tatelkuz Lake.  Site runoff water will be stored on site within the TSF, and the supernatant will be recycled back to the process plant.

The design of the TSF and its location near the open pit will minimize the Project footprint, prevent surface effluent discharge during operations, and allow for simple and effective water management at closure.

1.18 Infrastructure

The planned on-site infrastructure for the Project includes the following:

· Electrical substation

· Mine haul roads and access roads for plant and site facilities

· Explosives plant/magazine

· Mine truckshop and associated warehouse

· Diesel fuel storage tanks and fuel station

· Site offices

· Warehouse and cold storage

· Medical and emergency response facilities

· Security gate

· Fresh water distribution, potable water treatment plant, and fire water distribution

· Construction laydown area

· Core logging area

· Microwave communications facility (existing)

· Camp and sewage treatment

· Helipad

The construction camp will be sized for up to 1,500 personnel and will be modified and refitted into the operations camp for approximately 400 personnel.  Employees are expected to live in several local and regional communities.  Bus transportation to site will be provided for employees.

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1.19 Water Management

The water management strategy is to utilize water within the Project area to the maximum practical extent.  The plan involves collecting and managing site run-off from disturbed areas and maximizing the recycle of process water.  Site runoff water will be stored on site within the TSF and used to inundate the potentially acid generating (PAG) tailings solids and waste rock.  Excess water will be stored in the supernatant pond within the TSF and recycled to the mill for use in the process.

The key facilities for the water management plan are:

· Open pit

· Mill (including fresh and process water tanks)

· Tailings storage facility (TSF)

· Diversion and water management structures

· Fresh water supply

· Sediment and erosion control measures for the facilities

Sediment and erosion control strategies will include establishing diversion and run-off collection ditches, constructing sediment control ponds, and stabilizing disturbed land surfaces to minimize erosion.

1.20 Power and Electrical

The PEA assumes construction of a 133 km long, 230 kV transmission line to supply the Project with power.  The line will have an interconnection to the BC Hydro system at the GLN Substation near Fraser Lake.  The transmission line corridor will run east from the GLN Substation for approximately 19 km and continue south for another 114 km, following existing logging roads to the Project.

Future planned studies will include collection of detailed environmental baseline data, completion of a BC Hydro System Impact Study, First Nations and public consultation, and an in-depth constructability field review.

1.21 Water Supply

The preliminary water balance for the Project indicates a water deficit during operations.  Water will be required to mitigate flow reductions in Davidson Creek and sustain fisheries.  Additional water will be required for the Project.

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A trade-off study was carried out to examine potential water supply sources from lakes within close proximity of the mine site.  The best alternative identified was Tatelkuz Lake, approximately 15 km northeast of the Project site because the lake has the largest catchment and inflow, and the projected withdrawal rate would result in the least effect on flows, lake volumes, and associated environmental conditions.

A 15.7 km long pipeline consisting of 18 inch diameter steel pipe is proposed.  Two stages of pumping will be required to maintain design pressure along the line because of the 416 m increase in elevation between the lake and the discharge point (TSF) at the Project site.

1.22 Capital Costs

The total estimated cost to design, construct, and commission the 60,000 t/d gold facilities described for the Project is $1,814 million.  A parity USD:CAD exchange rate has been assumed for the purpose of determining the capital cost in US$, therefore a capital cost of US$1,814 million is assumed.

The capital cost estimate for the Project covers the costs to design, procure, construct, and commission the facilities described in this report.  The estimate is categorized as an AMEC Class 5 Scoping Level Estimate, in Q2 2012 Canadian dollars, with an expected accuracy of ±25%.  The estimate was based on a combination of semi-detail estimating for some components of the Project and capacity factor or equipment factor estimating for others.  Capital costs for surface facilities include the purchase of materials and equipment, construction and installation of all structures, utilities, materials, and equipment, and all associated indirect and management costs.  They also include contractor and engineering support to commission the process plant to ensure all systems are operational.  The estimate excludes off-site facilities, such as an off-site reagent transfer and storage facility.

Capital costs are summarized in Table 1-3.

It is assumed that $77 million salvage value can be obtained from major mine equipment and the crushing and grinding equipment from the processing plant.

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Table 1-3: Initial Capital Cost Summary by Major Area

Area	Description	Cost ($M)
1000	Mining	208
2000	On-Site Infrastructure	181
3000	Processing Plant	539
4000	Water/Waste Management	74
5000	Project Access Corridor	16
7000	Off-Site Facilities	69
	Subtotal Direct Cost	1,087
9000	Indirect Cost	327
9100	Owner’s Cost	54
	Subtotal Indirect & Owner’s Costs	381
	Total Direct, Indirect & Owner’s Costs	1,468
PP00	Contingency	346
	Total Project Cost	1,814

1.23 Operating Costs

The average life-of-mine (LOM) operating cost including royalty for Blackwater is estimated at $15.17/t milled, and $13.01/t after silver credits (Table 1-4).  After accounting for a silver credit of US$90/oz Au produced, and the royalty payments, the average LOM operating cost is US$543/oz Au produced.  Table 1-5 shows the cost by periods during the mining operation and Figure 1-1 shows the annual gold production and cash costs.

A total labour loading factor of 32% has been applied to labour costs to allow for benefits including vacation, retirement savings, pension plan contributions, unemployment insurance, workers compensation, sick leave, and training.  Consistent energy cost assumptions have been applied to mining, processing, and general and administration costs.

A diesel cost of $1.04/L has been assumed, based on an average WTI crude oil price of US$90/bbl and exchange rate of CAD$1 = US$0.94.  An average electricity tariff rate of 4.44 c/kWh was assumed for Blackwater, based on the BC Hydro tariff rate effective at May 2011, Schedule 1823: Stepped Rate, BC Zone 1 (BC Hydro, 2012).  Power consumption has been estimated for the mining, processing, and G & A areas, and power costs have been allocated proportionally.

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Table 1-4: Summary of Blackwater LOM Operating Costs

	LOM Unit Cost
Area	($/t milled)	($/oz Au produced)	(US$/oz Au produced)
Mining	6.21	276	259
Processing	7.59	337	317
G&A	0.95	42	40
Refining Costs	0.23	10	9
Royalties (0.56%)	0.18	8	8
Total Cash Costs	15.17	674	633
Silver Credits	(2.16)	(96)	(90)
Total Cash Costs incl. silver credit and royalties	13.01	578	543

Table 1-5: Cash Costs per Ounce of Gold Produced During Periods of Mine Operation

		Years
		1 – 5	1 – 15	16 - 17	LOM
Cash Costs - $/oz inc. royalty		579	666	811	674
Cash Costs - $/oz inc. Ag credit, royalty		497	570	722	578
Cash Costs - US$/oz inc. royalty		544	626	762	633
Cash Costs - US$/oz inc. Ag credit, royalty		467	536	678	543

Figure 1-1: Gold Production and Cash Costs (net of royalty and silver credits)

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1.24 Markets

New Gold has an operative refining agreement with Johnson Matthey in Salt Lake City for refining of gold and silver doré produced from New Gold’s Mesquite Mine.  New Gold’s share of the bullion is sold on the spot market, by marketing experts retained in-house by New Gold.  New Gold intends to market doré from Blackwater in a similar manner to that used for Mesquite.  The terms contained within the sales contracts are typical of, and consistent with, standard industry practice, and are similar to contracts for the supply of doré elsewhere in the world.

In the opinion of the QPs, New Gold will be able to market gold produced from the Project.  Sales contracts that could be negotiated would be expected to be within industry norms.  However, the majority of production would be expected to be spot marketed.

1.25 Economic Analysis

The following section is preliminary in nature and partly based on Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the PEA based on these Mineral Resources will be realized.  Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.  Dates cited are for illustrative purposes only, as a decision to proceed with mine construction will require formal approvals from appropriate regulatory authorities and from New Gold’s Board, and and will require support from additional, more detailed, studies, and declaration of Mineral Reserves.

The results of the economic analyses discussed in this section represent forward-looking information as defined under Canadian securities law.  The results depend on inputs that are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here.  Information that is forward-looking includes:

· Mineral Resource estimates

· Assumed commodity prices and exchange rates

· The proposed mine production plan

· Projected recovery rates

· Infrastructure construction costs and proposed operating costs

· Project development schedules

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

· Assumptions that an EA will be approved by Provincial and Federal authorities.

· Assumptions that New Gold will be able to acquire the social licence to operate

A discounted cash flow (DCF) model was developed to assess the value of the Project throughout the PEA phase, where different project configurations were evaluated with the objective of selecting a base case.  The DCF model developed assumes 100% equity financing (no debt), and all figures are expressed in Q2 2012 constant dollars or real terms.  No inflation or real cost escalation was assumed.  Cash flows are assumed to occur mid-period.  The Project value is determined on a pre-tax basis and discounting was performed to the start of 2015, which marks the first year of proposed project construction.  Therefore, expenditures incurred prior to 2015 are considered sunk and are not considered in the analysis.  These expenditures cover additional exploration and project studies.  Mining pre-production is planned to start in 2016 and 2017 would mark the first year of revenue.

The DCF model is based on the underlying economic assumptions listed in Table 1-6.

Table 1-6: Economic Inputs

Metal Price
Gold	US$/oz	1,275
Silver	US$/oz	22.5
Discount rate	%	5.0
Exchange rate (1 CAD = X USD)	CAD/USD	0.94
Refinery Deductions
Payable factor – Au	%	99.9
Payable factor – Ag	%	95.0
Dore refining treatment charge	$/oz Au + Ag	0.80
Au transport and insurance	$/oz Au	0.32

The economic model includes royalties payable on Claims 515809 and 515810.  It also assumes that the option to buy out the royalty on claims 509273, 509274, and 509275 for $2 M will be exercised in 2016, as this buyout price is substantially less than the NSR royalty on these claims.  Over the LOM, mining costs are $1.90/t moved.  Processing costs average $7.59/t milled.  General and administrative costs are variable and are estimated at $0.95/t milled.  The initial capital cost is $1,814 million.  Sustaining capital will be split among three areas:  TSF, mining, and plant.  Sustaining capital will total $571.5 M and average $33.6 M/a.

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The calculated pre-tax NPV5 for the base case is $1,872 million. After conversion to US$, the pre-tax NPV5 for the base case is US$1,655 million, the IRR is 16.4% and the simple payback (undiscounted) period is 4.7 years from the start of production.  Allowing for taxes, the post-tax NPV5 for the base case is US$1,123 million, the IRR is 14.0% and the simple payback period is 4.8 years.

Average life of mine cash costs, net of silver credits and royalties, are US$543/oz.

Average annual metal production over the initial 15 years of mine life is forecast to be 507,000 ounces (oz) Au and 2,039,000 oz Ag.  There will be 1.4 subsequent years of production where lower-grade stockpiles are processed.  The mine plan has been designed for extraction of higher gold grade resources in the early years.  Over the first five years of operation, the average gold feed grade will be 0.95 g/t and average annual gold production will be 569,000 oz per year.  Total LOM gold and silver production including lower-grade stockpiles will be 6.2 million oz and 18.6 million oz from the Indicated category and 1.8 million oz and 13.5 million oz from the Inferred category respectively.  Based on metal prices of US$1,275/oz Au, US$22.50/oz Ag and an exchange rate of 0.94 (US$:CAD$), gold will account for 93% of the revenue and silver will account for 7%.

A summary of the cashflow analysis is presented in Table 1-7.

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Table 1-7: Production Summary

	Units	LOM Total	LOM Annual Average	Year 1–5 Annual Average	Year 1–15 Annual Average	Year 16–17 Annual Average
Life of Mine1	Years	16.4	NA
Metal Price
Gold	US$/oz	NA	1,275
Silver	US$/oz	NA	22.5
Production
Mineralized Material Milled - Indicated	kt	255,935	15,606	17,239	15,764	13,913
Mineralized Material Milled - Inferred	kt	99,992	6,097	4,223	5,990	7,242
Waste Mined	kt	839,102	51,165	61,040	59,073	14,157
Strip Ratio W:MM	NA	2.36		2.42	2.72	0.66
Gold grade - Indicated	g/t	0.87		1.00	0.90	0.49
Gold grade - Inferred	g/t	4.3		5.4	4.4	3.3
Silver grade - Indicated	g/t	0.65		0.75	0.66	0.54
Silver grade - Inferred	g/t	7.9		4.7	8.5	3.1
Metal Feed
Gold - Indicated	koz	7,172
Gold - Inferred	koz	2,082
Silver - Indicated	koz	35,276
Silver - Inferred	koz	25,508
Process Recoveries
Gold	%	86.6
Silver	%	52.9
Metal Recovered
Gold - Indicated	koz	6,222
Gold - Inferred	koz	1,791
Silver - Indicated	koz	18,648
Silver - Inferred	koz	13,478
Gold - Annual production rate	koz		489	569	507	296
Silver - Annual production rate	koz		1,959	1,966	2,039	1,103
LOM Recovered Metal Value
Gold	M$	10,868
Silver	M$	769

1 Excludes pre-production year

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

1.26 Sensitivity Analysis

The Project NPV is most sensitive to changes in the gold price, then changes in the USD to CAD exchange rate, and less sensitive to changes in the operating cost (Figure 1-1).  It is least sensitive to capital cost changes.  The change in the gold price mirrors changes in deposit gold grades.  The sensitivity of NPV, IRR and payback period to a range of gold and silver prices, and exchange rate inputs is shown in Table 1-8.

Table 1-8: Sensitivity of Economic Performance vs. Metal Prices and Exchange Rates (base case is highlighted)

Summary of Project Economics
Gold Price (US$ per ounce)	Silver Price (US$ per ounce)	US$/CDN$ foreign exchange	5% NPV (US$ billions)			IRR (%)			Payback Period (Years)
			Pre-Tax	After-tax		Pre-tax	After-tax		Pre-tax	After-tax
1,275	22.50	0.94	1,655	1,123		16.4	14.0		4.7	4.8
1,600	30.00	0.97	3,338	2,246		25.9	22.0		3.0	3.1
1,775	34.50	1.00	4,202	2,815		30.4	25.8		2.6	2.7
1,800	35.00	1.00	4,338	2,905		31.1	26.4		2.5	2.6

Figure 1-2: Selected Sensitivities: Major Drivers of Pre-Tax NPV (CDN$)

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1.27 Interpretation and Conclusions

The Project outlined using the assumptions in this Report has a positive economic return.  New Gold plans to commence more detailed studies in support of declaration of Mineral Reserves and detailed mine designs, including confirmation of the process flowsheet and waste material handling options.  New Gold intends to subsequently complete a feasibility study.

New Gold has performed a risk and opportunity analysis.  Risks identified include conversion of some or all of the Inferred mineral resources to higher-confidence mineral resource categories, process recoveries, amounts and exact leaching characteristics of waste materials, and cost escalations due to various factors.  Opportunities identified include delineation of additional mineralization that can support mineral resource estimates, steepening of pit walls if geotechnical conditions permit, and reductions in mining and processing costs.

1.28 Recommendations

AMEC recommends that New Gold considers the recommendations as activities which will support Project advancement should New Gold determine that a Feasibility Study (FS) is warranted.

The program is designed as a two-phase program, with all elements of the first phase of the program conducted concurrently; this work phase is in progress. Phase 1 work comprises various testwork programs. Phase 2 comprises a number of studies that will culminate in the completion of a FS, and will allow New Gold to submit an EA and EIS.

AMEC’s recommendations do not include provision for construction activities for site and access infrastructure such as the road.

The Phase 1 activities include infill and delineation drilling and assaying, data collection, metallurgical testing, engineering optimization studies and investigations and studies and activities to support the basis of the FS and public consultation processes. Some more specific recommendations for work focus have also been included for mineral resource estimation, road and powerline design, and plant design purposes. The total cost of these activities is estimated to be between about $90 M to $102 M.

The Phase 2 activity comprises completion of a Feasibility Study, an EA, an EIS, and on-going community and government engagement activities estimated at between about $38 M and $43 M.

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Total program costs are likely to range between approximately $128 M and $145 M.

AMEC notes that the Phase 1 work is already in progress, which includes infill, delineation and condemnation drilling and assaying, geotechnical drilling and testing for both the planned TSFs and the open pit, metallurgical sample collection, assaying and testing and discussions relating to road usage and grid power connection.

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2.0 INTRODUCTION

AMEC Americas Inc (AMEC) was requested by New Gold Inc. (New Gold) to prepare a National Instrument 43-101, Standards of Disclosure for Mineral Projects (NI 43-101) Technical Report (the Report) for the wholly-owned Blackwater gold project (the Project) located in central British Columbia, Canada (Figure 2-1).

2.1 Terms of Reference

The Report was prepared to support a preliminary economic analysis (PEA) on the Project, as disclosed in the New Gold news release entitled “New Gold Announces Positive Preliminary Economic Assessment at Blackwater Gold Project with Average Annual Gold Production over 500,000 Ounces at Well Below Industry Average Cash Costs” and dated September 20, 2012.

2.2 Qualified Persons

The following served as Qualified Persons (QPs) as defined in NI 43-101:

· Ronald G. Simpson, P Geo. (GeoSim Services Inc.)

· Herbert E. Welhener, MMSA - QPM (Independent Mining Consultants Inc.)

· Bruno Borntraeger, P. Eng. (Knight Piésold Ltd.)

· Ignacy (Tony) Lipiec, P. Eng. (AMEC)

· Ramon Mendoza Reyes, P Eng. (AMEC)

2.3 Site Visits and Scope of Personal Inspection

Mr. Simpson visited the site from on December 13, 2010; September 8, 2011; November 28; 2011, and September 20, 2012.  The purpose of the visits was to review the drilling, sampling, and quality assurance/quality control procedures.  The geology and mineralization encountered in the drill holes completed to date were also reviewed.  A detailed description of the site visit findings is included in Section 12.1.

Mr. Borntraeger visited the site on September 28, 2011; December 20, 2011; February 21, 2012; and April 24, 2012.  The purpose of the visits was to review the drilling, sampling, and quality assurance/quality control procedures for the geotechnical program.

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2.4 Effective Dates

There are a number of effective dates, as follows:

· Date of the last assay data in the database supporting Mineral Resource estimation:  May 14, 2012

· Date of the closeout of the database supporting Mineral Resource estimation:  June 8, 2012

· Date of the Mineral Resource estimate:  July 18, 2012

· Date of the supply of the last information on mineral tenure:  September 5, 2012

· Date of the economic analysis used in the PEA:  August 28, 2012

· Date of the last supply of information on the ongoing drilling program:  August 28, 2012

The overall effective date of the Report is August 28, 2012 which is the date of the supply of the last drill information from the Project, and the date of the economic analysis in the PEA.  Drilling is ongoing on the Project.

2.5 Information Sources and References

Reports and documents listed in the Reliance on Other Experts (Section 3.0) and References (Section 27.0) sections of this Report were used to support the preparation of the Report.

2.6 Previous Technical Reports

New Gold has previously filed technical reports on the Project as follows:

· Simpson, R., 2011a:  Technical Report, Blackwater Gold Project, Omineca Mining Division, British Columbia, Canada:  report prepared for New Gold Inc. and Silver Quest Resources Ltd., effective date March 2, 2011, re-addressed June 6, 2011.

· Simpson, R., 2011b:  Technical Report, Blackwater Gold Project, Omineca Mining Division, British Columbia, Canada:  report prepared for New Gold Inc. and Silver Quest Resources Ltd., effective date September 19, 2011.

· Simpson, R., 2012:  Technical Report, Blackwater Gold Project, Omineca Mining Division, British Columbia, Canada:  report prepared for New Gold Inc. and Silver Quest Resources Ltd., effective date March 7, 2012.

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3.0 RELIANCE ON OTHER EXPERTS

The QP authors of this Report state that they are qualified persons for those areas as identified in the "Certificate of Qualified Person" for each QP, as included in this Report.  The QPs have relied, and believe there is a reasonable basis for this reliance, upon the following other expert reports, which provided information regarding mineral rights, surface rights, and environmental status in sections of this Report as noted below.

3.1 Mineral Tenure

The QPs have not reviewed the mineral tenure, nor independently verified the legal status, ownership of the Project area, underlying property agreements or permits.  The QPs have fully relied upon, and disclaim responsibility for, information derived from legal experts and from New Gold for this information through the following documents:

· Lawson Lundell LLP:  New Gold Inc.:  opinion letter on corporate status of New Gold, addressed to New Gold from Lawson Lundell LLP dated September 19, 2012.

· Lawson Lundell LLP:  New Gold Report Summarizing Certain Agreements Pertaining to the Blackwater Project dated September 19, 2012.

· Lawson Lundell LLP:  New Gold Inc. Blackwater Project:  opinion letter addressed to New Gold from Lawson Lundell LLP dated September 19, 2012, regarding the recorded holdings of the mineral claims comprising the Project.

This information is used in Sections 4.3 and 4.5 of the Report.

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3.2 Surface Rights

The QPs have fully relied upon, and disclaim responsibility for, information supplied by experts retained by New Gold for information relating to the status of the current Surface Rights as follows:

· Lawson Lundell LLP:  New Gold Report on Certain Surface Rights in the Vicinity of the Blackwater Project dated September 19, 2012.

This information is used in Section 4.4 of the Report.

3.3 Taxation

The QPs have fully relied upon, and disclaim responsibility for, information supplied by New Gold experts for information relating to the taxation assumptions used in the after-tax financial model as follows:

· Letter from New Gold dated October 10, 2012.

This information is used in Section 22.4.

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4.0 PROPERTY DESCRIPTION AND LOCATION

The Blackwater Gold Project lies in central British Columbia, approximately 112 km southwest of Vanderhoof and 446 km northeast of Vancouver.  The property is centred at 5893000 N and 375400 E and is located in NTS sheet 93F/02 (Figure 4-1 and Figure 4-2).

4.1 Property and Title in British Columbia

4.1.1 Mineral Title

Prior to June 1, 1991, mineral tenure was manually recorded on, or attached to, the original application document for a mineral claim or the original lease document for a mining lease.  From June 1, 1991 to January 11, 2005, all records were entered into a computer database, maintained by the Gold Commissioner’s Office.  On January 12, 2005, the British Columbia mineral titles system was converted to an online registry system, MTO, and ground-staking of claims were eliminated in favour of map-staking based on grid cells.

Claims lodged prior to January 12, 2005 are referred to as legacy claims.  At the date of changeover to map-staking, claim holders could convert legacy claims to cell claims, or maintain the original legacy claim.

Mineral title may also be held as part of Crown grants or freehold tenure issued under separate grant, such as a railway grant.  Crown-granted mineral rights originate from staked mineral claims that were surveyed then granted from the Crown to private individuals or corporations under the legislation in effect at the time of grant.  There can be instances where there may be more than one type of mineral tenure in existence over the same land area; examples are where a Crown-granted mineral title is overlapped by a title granted under the MTO.  In this case, the holder of the MTO grant is entitled only to those minerals not covered in the Crown-granted mineral title.

To keep claims in good standing in accordance with the Mineral Tenure Act, a minimum value of work or cash-in-lieu is required annually.  The value of exploration and development required to maintain a mineral claim for one year is currently set at $5 per hectare for the first and second anniversary years, $10 per hectare for the third and fourth anniversary years, $15 per hectare in the fifth and sixth anniversary years, and $20 per hectare for each subsequent anniversary year.  The required payment to maintain a mineral claim for an anniversary year is double the value of exploration and development that would be required to maintain the claim.

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The holder of a mineral claim or mining lease issued under the Mineral Tenure Act does not have exclusive possession of the surface or exclusive right to use the surface of the land.  However, the holder of such claims and leases does have the right to access the lands for the purpose of exploring for minerals and to use the surface for mining activities (exploration, development, and production).

The surface of a mineral claim or mining lease may either be privately owned or owned by the Crown.

4.2 Tenure History

The Project area was initially explored by Granges Inc from 1973.  In 2005, Silver Quest Resources Ltd. (Silver Quest) acquired an interest in the Project, and Richfield Ventures Corp. (Richfield) joint ventured into the Project in 2009.

In 2011, New Gold acquired Richfield and Silver Quest, and a third company, Geo Minerals Limited (Geo), to consolidate the ground holdings in the Project area.

4.3 Mineral Tenure

The Project consists of 69 mineral claims, covering an area of 27,175 ha (Figure 4-2).  A total of 66 of the claims are recorded in the name of New Gold.  The remaining three claims, 637203, 637205, and 637206, termed the JR Claims, are recorded in the name of third parties.  New Gold was granted an option to acquire a 100% interest in the JR Claims, subject to a royalty.  No opinion was provided as to unrecorded interests in or affecting the claims as these cannot be ascertained through inquiries with the Gold Commissioner’s office.  The mineral claims are summarized in Table 4-1.

New Gold has moved to have all claims with a common expiry date as indicated in Table 4-1.

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Table 4-1: Mineral Tenure Summary Table

Tenure Number	Claim Name	Issue Date	Good To Date	Status	Area (ha)
503050	WHITEWATER	13/01/2005	29/08/2022	GOOD	348.76
509273	GOT	19/03/2005	29/08/2022	GOOD	484.45
509274	got2	19/03/2005	29/08/2022	GOOD	38.75
509275	got3	19/03/2005	29/08/2022	GOOD	19.38
515809*		07/01/2005	29/08/2022	GOOD	581.6
515810*		07/01/2005	29/08/2022	GOOD	348.96
536650	NIGHT FLIGHT	07/06/2006	29/08/2022	GOOD	271.42
602167	BWD	04/05/2009	29/08/2022	GOOD	387.94
602168	BWD2	04/05/2009	29/08/2022	GOOD	310.35
607194	BLACKWATER 2	07/08/2009	29/08/2022	GOOD	464.86
607195	BLACKWATER 1	07/08/2009	29/08/2022	GOOD	348.74
630903	BW1	09/09/2009	29/08/2022	GOOD	465.33
630944	BW2	09/09/2009	29/08/2022	GOOD	251.9
630963	BW3	09/09/2009	29/08/2022	GOOD	465.15
630983	BW4	09/09/2009	29/08/2022	GOOD	387.29
630984	BW5	09/09/2009	29/08/2022	GOOD	464.98
631003	BW6	09/09/2009	29/08/2022	GOOD	484.14
631024	BW7	09/09/2009	29/08/2022	GOOD	445.49
631043	BW8	09/09/2009	29/08/2022	GOOD	464.73
636583	KASSY 1	18/09/2009	29/08/2022	GOOD	464.53
636603	KASSY 2	18/09/2009	29/08/2022	GOOD	464.27
636604	KASSY 3	18/09/2009	29/08/2022	GOOD	464.01
636623	KASSY 4	18/09/2009	29/08/2022	GOOD	463.82
636643	KASSY 5	18/09/2009	29/08/2022	GOOD	483.3
636644	KASSY 6	18/09/2009	29/08/2022	GOOD	483.47
636663	KASSY 7	18/09/2009	29/08/2022	GOOD	290.18
636683	RIGHT STUFF 1	18/09/2009	29/08/2022	GOOD	464.86
636684	RIGHT STUFF 2	18/09/2009	29/08/2022	GOOD	464.87
636703	RIGHT STUFF 3	18/09/2009	29/08/2022	GOOD	464.87
636723	RIGHT STUFF 4	18/09/2009	29/08/2022	GOOD	464.45
636724	RIGHT STUFF	18/09/2009	29/08/2022	GOOD	464.63
636725	RIGHT STUFF 6	18/09/2009	29/08/2022	GOOD	484.18
636727	RIGHT STUFF 7	18/09/2009	29/08/2022	GOOD	484.41
636743	RIGHT STUFF 8	18/09/2009	29/08/2022	GOOD	483.84
636763	RIGHT STUFF 9	18/09/2009	29/08/2022	GOOD	464.85
636764	RIGHT STUFF 10	18/09/2009	29/08/2022	GOOD	484.56
636765	RIGHT STUFF 11	18/09/2009	29/08/2022	GOOD	465.12
636766	RIGHT STUFF 12	18/09/2009	29/08/2022	GOOD	464.76
636767	RIGHT STUFF 13	18/09/2009	29/08/2022	GOOD	464.42

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Tenure Number	Claim Name	Issue Date	Good To Date	Status	Area (ha)
637203#	Ozzy	19/09/2009	29/08/2022	GOOD	484.42
637205#	Baby Jane	19/09/2009	29/08/2022	GOOD	464.57
637206#	David Dale	19/09/2009	29/08/2022	GOOD	464.57
640804	PUREANDY	25/09/2009	29/08/2022	GOOD	310.16
641685	RICHFIELDADJACENTCC	26/09/2009	29/08/2022	GOOD	445.58
642043	BW	27/09/2009	29/08/2022	GOOD	232.57
642063	BW 2	27/09/2009	29/08/2022	GOOD	232.66
642064	BW3	27/09/2009	29/08/2022	GOOD	310.34
834367	RICH 1	27/09/2010	29/08/2022	GOOD	484.06
834371	DAVIDSON	27/09/2010	29/08/2022	GOOD	425.87
834533	DAVIDSON 1	29/09/2010	29/08/2022	GOOD	77.39
834534	DAVIDSON 2	29/09/2010	29/08/2022	GOOD	406.33
834923	DAVIDSON 3	10/02/2010	29/08/2022	GOOD	483.59
834924	DAVIDSON 4	10/02/2010	29/08/2022	GOOD	483.45
834926	DAVIDSON 5	10/02/2010	29/08/2022	GOOD	483.45
834948		10/03/2010	29/08/2022	GOOD	484.69
834998	RICH 2	10/04/2010	29/08/2022	GOOD	426.28
835005		10/04/2010	29/08/2022	GOOD	465.50
835009		10/04/2010	29/08/2022	GOOD	271.27
835011		10/04/2010	29/08/2022	GOOD	484.69
835012		10/04/2010	29/08/2022	GOOD	484.46
835013		10/04/2010	29/08/2022	GOOD	174.41
835014	DAVE	10/04/2010	29/08/2022	GOOD	116.23
835016		10/04/2010	29/08/2022	GOOD	232.77
835019		10/04/2010	29/08/2022	GOOD	387.77
835020		10/04/2010	29/08/2022	GOOD	329.47
835021	BW WEST	10/04/2010	29/08/2022	GOOD	387.94
835022	BW WEST2	10/04/2010	29/08/2022	GOOD	368.56
835023		10/04/2010	29/08/2022	GOOD	465.19
835025	BW WEST2	10/04/2010	29/08/2022	GOOD	38.79

Note:

* Claim 515809 is referred to as the Dave Claim and Claim 515810 is referred as the Jarrit Claim

# denotes the JR Claims (recorded in the names of third-party optionors).

The currently-defined Blackwater deposit spans one of the Davidson claims (509273, but not 509274, and 509275), the Dave claim (515809) and the Jarrit claim (515810).

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

4.4 Surface Rights

The claims are located on Crown land.  Limited searches that have been conducted did not reveal any dispositions of surface rights by the Crown under the Land Act within the Project area, although surface use permits, licences and authorizations have been granted.  Table 4-2 summarizes the active government reserves, grazing licences, forest licences to cut, special use permits, outfitter and trapline permits in the Project area.

Table 4-2: Surface Use Permits

Tenure Number	Permit Details
636727	Active Land Act reserve No 245442, partially overlaps; held by Ministry of Forests, Lands and Natural Resource Operations as a UREP/Recreation reserve
602168	Active Land Act reserve No 250098 partially overlaps; held by Ministry of Forests and Range as a communications reserve/notation
835022	Active grazing licence No 1604762 partially overlaps
636604, 636623, 636643, 636644, 636767, 834924	Active grazing licence No 185651 overlaps or partially overlaps
636604, 636623, 834923, 834924, 834926	Active grazing licence No. 1431583 partially overlaps
636723	Active forest recreation tenure No 235653 partially overlaps
636604, 636603, 834924	Active forest licence No 275265 partially overlaps
834926	Active forest licence No. 262858 partially overlaps
607195	Active forest licence No 1280949 partially overlaps
835016, 835005	Active forest licences No 1803965 and 1803963 partially overlap
835016	Active forest licence No 1803966 partially overlaps
835022	Active forest licences No 1804968 and 1804961 partially overlap
503050, 631003, 631024, 631043, 630963, 630984	Active Occupant Licence to Cut No 1806968 overlaps or partially overlaps
503050, 509273, 509274, 509275, 515809, 607194, 630944, 630983,631024,631043,637206, 640804, 641685, 834367, 834371, 515810, 630963, 630984, 631003, 636583, 642043	Active Occupant Licences to Cut No 1778742 overlaps or partially overlaps
503050, 509273, 509274, 509275, 515809, 607194, 630944, 630983, 631024, 631043, 637206, 640804, 641685, 834367, 834371	Active Occupant Licences to Cut No 1778744 overlaps or partially overlaps
503050, 509273, 515809, and 515810	Active Occupant Licence to Cut No 1778743 partially overlaps
641685, 642043, 607194	Active Occupant Licence to Cut No. 1806907 partially overlaps

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Tenure Number	Permit Details
503050, 509273, 607194, 607195, 631024, 835009, 835014	Active Special Use Permit No. 1606182 partially overlaps
636583, 636603, 636604, 636623, 636643, 636644, 636663, 636723, 636724, 636725, 636743, 636763, 636766, 636767, 637205, 637206, 834367, 834371, 834533, 834534, 834923, 834924, 834926	Active trapline TR0601T003 overlaps or partially overlaps
503050, 509273, 509274, 509275, 515809, 515810, 536650, 602168, 607194, 630903, 630944, 630963, 630983, 630984,631003, 631024, 631043, 636583, 636683, 636684, 636703, 636725, 636727, 636763, 636764, 636765, 636766, 637203, 637205, 637206, 640804, 641685, 642043, 642063, 642064, 834367, 834371, 834948, 834998, 835005, 835011	Active trapline TR0512T027 overlaps or partially overlaps
515809, 515810, 536650, 602167, 602168, 630903, 630963	Active trapline TR 0512T014 overlaps or partially overlaps
607195, 835019, 835020, 835021, 835022, 835023, 835025, 835013	Active trapline TR 0601T002 overlaps or partially overlaps
515809, 536650, 602168, 607194, 607195, 630944, 630983, 631024, 631043, 636583, 636603, 636604, 636623, 636643, 636644, 636663, 637206, 640804, 641685, 642043, 642063, 642064, 834367, 834371, 834923, 834924, 834926, 834948,834998,835005, 835009, 835011, 835012, 835013, 835014, 835016, 835019, 835020, 835021, 835022, 835023, 835025	Active guide outfitter area certificate no. 600384 overlaps or partially overlaps
515809, 515810, 602167, 602168, 630903	Active guide outfitter area certificate no. 500929 overlaps or partially overlaps
503050, 509273, 509274, 509275, 515809, 515810, 536650, 607194, 630903, 630944, 630963, 630984, 631003, 631024, 631043, 636583, 636643, 636644, 636663, 636683, 636684, 636703, 636723, 636724, 636725, 636727, 636743, 636763,636764, 636765, 636766, 636767, 637203, 637205, 637206, 640804, 641685, 642043, 834533, 834534	Active guide outfitter area certificate no. 601039 overlaps or partially overlaps

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

4.5 Royalties and Encumbrances

Royalties on the claims are summarized in Table 4-3.  No other material encumbrances against the recorded holdings of the claims were identified in the legal opinions and reports provided to AMEC or in the information provided by New Gold, when all reports were read together.

Table 4-3: Details of Royalties over Blackwater Project Claims

Claims	Royalty Payable	Option to Purchase Royalty	Remaining Royalty Payable after Purchase Option is Exercised	Location	Comment
509273, 509274, and 509275	2% NSR	Option to purchase full NSR for $2 M	Nil	Claim 509273 covers a portion of the Blackwater deposit.  Remaining two claims are outside Blackwater deposit area	Option to purchase has not been exercised as at the date of this report.  The financial analysis in Section 22 assumes the option to purchase the 2% NSR will be exercised.
515809	2.5% NSR	Option to purchase 1% NSR for $ 1 M	1.5% NSR	Claim covers a portion of the Blackwater deposit.	Option to purchase the 1% NSR exercised.
515810	2% NSR	Option to purchase 1% NSR for $ 1.2 M	1% NSR	Claim covers a portion of the Blackwater deposit.	Option to purchase the 1% NSR exercised.
637203, 637205, and 637206	3% NSR	Option to purchase 2% NSR for $ 1 M (at any time after New Gold exercises its option to acquire a 100% interest in the claims)	1% NSR	Outside Blackwater deposit area	Option to purchase has not been exercised
835014	2% NSR	Option to purchase 1% NSR for $ 1 M	1 % NSR	Outside Blackwater deposit area	Option to purchase has not been exercised

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

4.5.1 Davidson Claims

Mineral claims 509273, 509274, and 509275 were acquired by Southern Rio Resources Ltd. (which later changed its name to Silver Quest), from West Range Exploration Ltd.  The vendor retained a 2% net smelter return royalty (NSR), which Southern Rio had an option to purchase for $2,000,000 at any time.  The option to purchase the royalty transferred to New Gold with the acquisition of Silver Quest and subsequent amalgamation of Silver Quest into New Gold.  Claim 509273 covers a portion of the Blackwater deposit.

4.5.2 Dave Option

New Gold, through Richfield, acquired a 100% interest in mineral claim 515809 (Dave Option) in 2010 from Jane Roderick, Rebekah Antkow, David Rozek, Benjamin Rozek, and John Blackwell.  The optionors retained a 2.5% NSR.  New Gold has purchased 1% of the Dave NSR royalty for $1,000,000, and a 1.5% NSR royalty remains.  The claim covers a portion of the Blackwater deposit.

4.5.3 Jarrit Option

New Gold, through Richfield, acquired a 100% interest in mineral claim 515810 (Jarrit Option) in 2010 from Jane Roderick, Rebekah Antkow, David Rozek, and Benjamin Rozek.  The optionors retained a 2% NSR royalty.  New Gold has purchased 1% point of the Jarrit NSR royalty for $1,200,000, and a 1% NSR royalty remains.  The claim covers a portion of the Blackwater deposit.

4.5.4 JR Option

In January 2011, Richfield acquired an option to earn a 100% interest in the JR Claims, 637203, 637205, and 637206, from the same optionors as the Jarrit Option.  The option has not yet been exercised by New Gold.  The optionors will retain a 3% NSR royalty upon exercise of the option by New Gold.  New Gold may purchase 2% points of the JR Claims NSR royalty for $1,000,000 at any time after it exercises of the option to acquire a 100% interest in those claims.

4.5.5 Saulnier Claim

New Gold, through Geo, acquired a 100% interest in mineral claim 835014 from Paul Saulnier.  Saulnier retained a 2% NSR royalty.  New Gold may purchase a portion of the NSR royalty for $1,000,000, such that a 1% NSR royalty will remain.

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

4.6 Permits

Permits required for Project development activities are discussed in Section 20.0.

4.7 Environmental Liabilities

Environmental liabilities associated with the Project are discussed in Section 20.0.

4.8 Social License

Socio-economic issues for the Project are discussed in Section 20.0.

4.9 Significant Risk Factors

Risk factors affecting the Project in terms of permit, environmental and social factors are outlined in Section 20.0.

4.10 Comments on Section 4

In the opinion of the QPs, the information discussed in this section supports the declaration of Mineral Resources and the preliminary economic assessment.  The QPs note the following:

· Information provided by legal experts and New Gold support New Gold’s ownership claims to the Project mineral tenure.

· Future development will require conversion of the mineral claims covering the Blackwater deposit to mining leases.

· The claims are located on Crown land.  Limited searches that have been conducted did not reveal any dispositions of surface rights by the Crown under the Land Act within the Project area, although surface use permits, licences and authorizations have been granted.  There is a reasonable expectation that with appropriate supporting studies as required by permitting authorities that land access and provision of land for infrastructure development for any proposed mining activity will be achievable.

· Five royalties are payable:  a 2% NSR on mineral claims 509273, 509274, and 509275; a 1.5% NSR on mineral claim 515809, a 1% NSR on mineral claim 515810, a 3% NSR on mineral claims 637203, 637205, and 637206, and a 2% NSR on mineral claim 835014.  However, the Blackwater deposit is affected by only three of these royalties:  the 2% NSR on mineral claims 509273, 509274, and 509275; the 1.5% NSR on mineral claim 515809, and the 1% NSR on mineral claim 515810.

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

· New Gold may purchase the 2% NSR on mineral claims 509273, 509274, and 509275 for $2 M, 2% NSR of the 3% NSR on mineral claims 637203, 637205, and 637206 for $1 M (at any time after it exercises the option to acquire a 100% interest in those claims), and 1% NSR of the 2% NSR on mineral claim 835014 for $1 M.

· For purposes of the economic analysis, it has been assumed that the % NSR on mineral claims 509273, 509274, and 509275 will have been acquired by New Gold and thereby extinguished, and that the only two remaining royalties that will affect the Blackwater deposit are the 1.5% NSR on mineral claim 515809, and the 1% NSR on mineral claim 515810.

Project No.:  169229
October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY

5.1 Accessibility

The Project site is readily accessible by vehicle from the Kluskus-Ootsa Forest Service Road originating approximately 10 km west of Vanderhoof, off Highway 16.  At kilometre 146.5 along this road, an 18 km mine road built in 1986 by Granges and improved by Richfield provides direct access to the property and camp location.  Driving time from Vanderhoof to the property is about 2.5 hours.

Helicopter access is from bases in Vanderhoof, Quesnel, or Prince George.

5.2 Climate

The climate in the Project area is sub-continental, characterized by brief warm summers and long cold winters resulting from the influence of cold arctic air.  The climate is also influenced by moisture-laden weather systems moving east by way of the low Kitimat Ranges.  Temperatures range from a minimum of -40°C in winter to a maximum of 32°C in summer.  The mean annual precipitation for the site is estimated to be 580 mm with 50% falling as rain and 50% as snow.

5.3 Local Resources and Infrastructure

The area is very sparsely inhabited; three ranches are found within a 20 km radius of the Project site.  Some services are available in Vanderhoof, but Prince George is the regional hub with air service from major centres.

There is no grid-connected power in the direct vicinity of the Project.  The main BC Hydro 500 kV transmission lines supplying western B.C. are approximately 100 km to the north.  Several interconnection points to existing 230 kV substations and transmission lines are possible in an area between Fraser Lake and Vanderhoof.  Power for the current Blackwater exploration camp is provided by generators.

The deposit is located on the north slope of Mt. Davidson, and the proposed project infrastructure will be sited predominantly in the Davidson Creek watershed.  Water for the Blackwater exploration camp is currently obtained from four groundwater wells that feed a potable water tank.

Infrastructure required for Project development is discussed in Section 18.0.

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

5.4 Physiography

The elevation of the Blackwater property ranges from just over 1,000 m in low-lying areas northeast of the proposed mine site to 1,800 m at the summit of Mt. Davidson on the southwest side of the property.

Outcrop on the property is limited.  Most of the area is covered with thick glacial deposits of 2 m or more, except for the upper 150 m of Mt. Davidson.  The slopes of Mt. Davidson are covered by till and colluvium, but the claim area is underlain by lodgement and melt-out till deposits.  Ice flow directions determined from glacial fluting are northeast.

The study area falls within the Fraser Plateau biogeoclimatic region and more specifically within the Nasko Upland subregion.  Five biogeoclimatic (BGC) units represent the study area, as follows:

· Sub-Boreal Spruce Moist Cold Babine Variant;

· Sub-Boreal Spruce Moist Cold Kluskus Variant;

· Engelmann Spruce – Subalpine Fir Moist Very Cold Nechako Variant;

· Engelmann Spruce – Subalpine Fir Moist Very Cold Parkland; and

· Boreal Altai Fescue Alpine Undifferentiated Subzone.

The Fraser Plateau biogeoclimatic region is a broad, rolling plateau of basaltic lava flows that was entirely overridden by ice sheets moving eastward in the north and northward in the south during the last glaciation.  The Fraser Plateau region contains ten subregions, with only the Nazko Upland occurring within the study area.

The Nazko Upland subregion is an undulating upland area that contains the Fawnie and Nechako ranges near the study area in the north-northeast.  Low-elevation valley bottoms are dominated by stands of lodgepole pine.  Hybrid white spruce tends to dominate on moist to wet sites below 1,500 m, while subalpine fir and Englemann spruce take over above 1,500 m.  Lodgepole pine is a major species on dry, fire-prone sites at most elevations.  The recent pine beetle epidemic has infested almost all of the lodgepole pine forests within this subregion.  The Nazko Upland subregion also contains an extensive network of lakes, rivers, and wetland complexes.  Atmospheric heating of these water bodies can result in convective activity and sporadic summer showers.

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

5.5 Regional Seismicity

The Project is situated more than 100 km west of the central region of the Coast Mountains in an area where the level of seismic activity has historically been low.  The maximum earthquake magnitude for the potential seismic source regions of northern B.C. and the Coast Mountains is estimated to be in the range of 7.0 to 7.3 (Adams and Halchuk, 2003).  There is also potential for smaller events in the interior of B.C.  Most recently, a magnitude 5.4 earthquake occurred near Prince George in 1986 but caused minor damage.

There has been much debate in recent years concerning the possibility of a large interplate earthquake of Magnitude 8 or 9 along the Cascadia subduction zone.  Such an event would likely take place at least 400 km southwest of the Project site.  The amplitude of ground motions experienced at the site would therefore be low due to attenuation over such a large distance.

5.6 Sufficiency of Surface Rights

There is sufficient suitable land area available within the mineral claims for any future tailings disposal, mine waste disposal, and installations such as a processing plant, and related mine infrastructure.

5.7 Comments on Section 5

In the opinion of the QPs:

· Mining activities should be capable of being conducted year-round.

· There is sufficient suitable land available for any future tailings disposal, mine waste disposal, and related mine infrastructure within the mineral claims.

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

6.0 HISTORY

Mineralization on the property was discovered in 1973 during a regional silt geochemical survey by Granges.  The survey located anomalous zinc and other metals in stream sediments east and north of the Project area.  Between 1973 and 1985 a range of ground and airborne geophysical and geochemical surveys was conducted to locate and delimit mineralization.

Drilling began in 1985 and continued to 1994 completing a total of 6,300 m in 75 holes.  A further 1,333 m of drilling in seven holes was completed in the winter of 2005-06.  The focus of the work during this period was on high-grade vein deposits, which at the time were considered the only known gold-silver targets in the Nechako Plateau.

Richfield began work on the Project in 2009 when the company recognized the bulk gold potential on the property.  In 2009, Richfield optioned the Davidson Claims from Silver Quest and the Dave Claim and Jarrit Claim from the Rozek family.  Richfield subsequently completed all of the earn-in requirements for these agreements, resulting in its holding a 75% interest in the Davidson Claims and 100% interests in each of the Dave and Jarrit Claims.

In June 2011, New Gold acquired all of the issued and outstanding common shares of Richfield pursuant to a court-approved plan of arrangement.  In December 2011, New Gold acquired all of the issued and outstanding common shares of Geo and Silver Quest.  New Gold subsequently amalgamated with Richfield, Geo and Silver Quest effective January 1, 2012 and became the direct operator of the Project.

Work completed by all operators is summarized in Table 6-1.  No production has occurred from the Project area.

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Table 6-1: Work History

Year	Operator	Work
1973	Granges	Regional silt survey located anomalous silver, zinc, and lead in the Mt. Davidson area.  This was followed by a wide-spaced soil survey northeast of Mt. Davidson.
1976	Granges	Soil sample and ground magnetometer surveys to follow up 1973 soil results.
1977	Granges	Pem claim staked covering most of the presently defined mineral deposit.  Pulse EM survey on the Pem claim (12.5 km)
1979	Granges	Vector Pulse EM survey on the Pem claim (7 km).
1981	Granges	Helicopter EM and magnetometer survey.
1981	Granges	Horizontal Loop EM survey on the Deb #1 claim.
1981	Granges	Reconnaissance mapping of the Mt. Davidson area.
1982	Granges	Soil geochemistry (220 samples) and ground magnetometer survey (20.8 line km) on the Pem claim.
1983	Granges	Hammer seismic survey.
1984	Granges	Hand-trenching (30 trenches for total of 66 m) and VLF survey (4.8 line km) on the Pem claim.  Only 1 trench intersected bedrock.
1985	Granges	Winkie drilling (8 holes for total of 507 m) on the Pem claim.  Holes DAV 1-8.
1986	Granges	Construction of access road from km 146.5 on the Kluskus Haulage road, east 18 km to the Pem grid.
1986	Granges	Percussion drilling (34 holes for  total of 1,524 m) on the Pem claim. RC 1-34.
1987	Granges	Diamond drilling (22 holes for total of 2,725 m) on the Pem claim. Holes DAV 9-31.
1992	Granges	Line Cutting (58.8km); collection of 955 soil samples and 35, stream silt samples; geological mapping over 6000 hectares at a scale of 1:10,000; 50 km of geophysical surveys ( IP, mag, VLF); , and diamond drilling of 5 holes BD92-32 to BD92-36, for total of 785 m.
1994	Granges	Linecutting (48.2 km); 29 rock, samples, 1598 soils samples,  23 silt samples and  4 lake sediments were collected;  Dighem airborne geophysical survey (881 line km of EM, Mag and radiometrics);  20km IP survey over the Dave claim; Diamond drilling (6 holes for total of 759 m). Holes DAV 94-37 to DAV 94-41.
1997	Kennecott Canada	4 km of Linecutting and an Induced Polarization survey on the Dave claim.
2005	Silver Quest	Diamond drilling (5 holes for total of 939 m). Holes DAV05-01 to DAV05-05.
2006	Silver Quest	Diamond drilling (2 holes for total of 394 m). Holes DAV06-06 and DAV06-07.
2007	Silver Quest	Soil sampling program, 335 samples.
2009	Richfield	Diamond drilling (18 holes for total of 3,665 m)  BW0042 to BW0059
2010	Richfield	Diamond drilling (57 holes for  total of 21,337 m). Holes BW0060 to BW0116.
2011	Richfield	Diamond drilling (59 holes for total of 19,734 m). Holes BW0117 to BW0175.
2011	Newgold	Diamond drilling (125 holes for total of 49,153 m). Holes BW0176 to BW0299. Additional 7 metallurgical test holes for a total of 2,282 m .
2012	Newgold	Diamond drilling (178 holes for total of 70,782 m). Holes BW0300 to BW0478.  Additional 5 geotechnical holes for a total of 1,800 m.

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

7.0 GEOLOGICAL SETTING AND MINERALIZATION

7.1 Regional Geology

The Project is underlain by rocks of the Stikine Terrane within an uplifted block termed the Nechako uplift.  The Nechako Plateau is an area of moderate relief between the Skeena Arch and the Stikine and Cache Creek Terrane contact.

The following discussion of regional geology is based on work by Diakow and Webster (1994), Diakow and Levson (1997), and Diakow and others (1997).  The southern Nechako Plateau near the Blackwater property is underlain by Late Triassic to Middle Jurassic island and continental arc assemblages and epicontinental sedimentary strata Near the Project, the volcaniclastic Hazelton Group represents the Stikine Terrane.  The Hazelton Group has a lower unit of felsic tuffs and sediments, named the Entiako Formation, and an upper unit of felsic to mafic volcanic flows and volcaniclastic and sedimentary rocks, termed the Naglico Formation.  Rocks of the Stikine Terrane in the southern Nechako Plateau are overlain by post-accretion Upper Jurassic volcaniclastic, sedimentary, and mafic to felsic volcanic rocks of the Bowser Lake Group.  The Hazelton and Bowser Lake groups in the area are collectively intruded by Late Cretaceous granitic to granodioritic plutons (e.g., Capoose batholith) exposed within the Nechako uplift.

Rocks of the Hazelton and Bowser Lake groups are overlain by Upper Cretaceous and Paleocene continental volcanic arc intermediate volcanic rocks and related sedimentary rocks of the Kasalka Group.  The Kasalka Group is a rock unit that occurs locally south of the Natalkuz fault.  Widespread Eocene volcanic arc-related extensional felsic volcanic rocks and minor sedimentary rocks of the Ootsa Lake Group overlie the older rocks and are themselves overlain on higher ridges by basalt and andesite of the Eocene Endako Group.

The Bowser Lake Group is poorly exposed in the southern Nechako Plateau, except within the Nechako uplift.  The Nechako uplift is a structurally raised block bounded by the Natalkuz fault to the north and the Blackwater fault to the south.  Both faults juxtapose older rocks of the Stikine Terrane (e.g., Hazelton and Bowser Lake groups) with younger Cretaceous and Eocene rocks of the Ootsa Lake and Endako groups.  This uplift provides a direct window through younger cover into older rocks.  The regional geologic setting is illustrated in Figure 7-1.

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Figure 7-1: Regional Geology

Note:  Figure courtesy New Gold, 2012

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7.2 Project Geology

Quaternary glacial overburden, colluvial, and fluvial deposits mask the majority of bedrock within the Project area.  Outcrop on the Blackwater property is sparse and limited to the peak and ridges of Mt. Davidson.  Property geology is based on historic and 2009/2010 drill core interpretations.  Figure 7-2 is a sketch map of the Project geology.

Interpreted geology is overlain on the regional geology as mapped by Diakow and Levson (1997).  The blue dashed line shown in Figure 7-2 is that of the 0.2 g/t Au grade shell contour from the mineral resource estimate discussed in Section 14.

The Project site is underlain by intercalated volcanic and volcaniclastic felsic to intermediate lapilli and ash tuff, volcanic breccias, and andesitic flows.  These strata form a local wedge of laterally discontinuous strata.  The Blackwater wedge is thought to dip generally northwest and is of limited aerial extent.  On the west the Blackwater wedge is faulted against younger massive felsic volcanic rocks of the Ootsa Lake Group.  The fault is a north-trending, presumed steep-dipping structure.  A similar relationship exists on the north side, where Blackwater host rocks are also juxtaposed next to Ootsa Lake Group strata across an east-northeast-trending fault.  Although displacement across the faults is not known, the relative age of rocks across the two faults implies that the Blackwater block is a horst or high-standing remnant and that the Ootsa Lake Group dropped to the west and north.

East and south of the Blackwater block the relationship to the country rocks is unknown but presumed to be stratigraphic.  It is likely the Blackwater strata are underlain by Bowser Lake Group beds.

The Blackwater wedge consists of a sequence of felsic to intermediate composition volcanic rocks.  Individual lithologies include felsic tuffs and lapilli tuffs, volcaniclasitic and epiclastic heterolithic breccias, and massive to layered andesites.  Dark reddish-brown anhedral equant, garnet crystal fragments up to a centimetre in diameter are common as an accessory in the heterolithic breccia and locally make up 1% to 2% of the rock.  XRF data on the garnets indicate they are Mn-rich spessartine.

Blackwater wedge rocks are pervasively hydrofractured and silicified.  The amount of silica introduced through hydrofracturing and silicification may amount to 25% or more of the total volume of volcanic rocks.  Although intensely hydrofractured, the Blackwater wedge lacks clearly recognizable large-scale faults or shear zones.  Instead, extensive zones of broken rocks are seen in the mineralized zone.  The zones grade laterally into unbroken rock and are generally not bounded by planar surfaces.

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Outcrops of massive felsic lapilli tuff assigned to the Ootsa Lake Group are found on the peak and ridges of Mt. Davidson.  This rock is generally darker grey than rocks drilled on the property, clasts are larger, and plagioclase porphyry dominates.  This unit is also characterized by fresh, black stubby euhedral doubly-terminated quartz crystals up to 1 mm across, which commonly make up a few percent of the rock.

At lower elevations near the Kluskus Road, bedded tuffs and sediments of the Jurassic Naglico Formation are observed.

The modified lithological codes used in the Blackwater drill database as of December 2011 are summarized in Table 7-1.

Table 7-1: Drill Database Lithological Codes

Code	Description
OB	Overburden
AND	Andesite
FT	Felsic tuff
FLPT	Fesic lapilli tuff
VC	Volcaniclastic
EC	Epiclastic
SED	Argillite / Sandstone / Conglomerate

7.3 Mineralization

Core drilling has defined a continuous zone of +0.2 g/t Au equivalent mineralization that extends at least 1,250 m along its longest dimension east-west and at least 1,000 m north-south.  The thickness of the zone ranges up to 600 m.

In general, all rocks at Blackwater are mineralized, whether this as trace pyrite-pyrrhotite-sphalerite in outboard andesite flows and volcaniclastics, or as gold-bearing polymetallic sulphide mineralization within the fragmental unit of the deposit.  The only exceptions are Eocene (?) dacite porphyry dykes intersected along the southern portion of the drilling grid, and amygdaloid mafic intermediate flows in the northern part of the grid, possibly related with the Eocene Ootsa Group.

Mineralized rocks within the main Blackwater resource area can be broadly divided into a thick succession of felsic to intermediate pyroclastic and volcaniclastic rocks, volcanic flows and breccias, and related volcanic and lithic-derived sedimentary units (fine to coarse epiclastic rocks).  Whole-rock analysis indicates that these units range from rhyolite to dacite to andesite in composition.  The age relationship between all of the intermediate and felsic volcanic units at Blackwater is not entirely understood, but the locally observed progressive inter-bedding of these units within the deposit implies a temporal relationship.

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Disseminated Au-Ag mineralization is defined by an east-west-trending tabular-conical shaped deposit with a lateral extent of up to 1,250 m east-west x 1,000 m north-south.  Mineralization remains open at depth in the southwestern portion of the deposit as well as to the north and northwest.  The centre of the deposit has an average thickness of 350 m with a vertical extension of up to 600 m where open.  The overall volume plunges shallowly to the north and northwest with inferred steep, north-plunging higher-grade mineralized shoots measuring tens of metres thick, likely influenced by near-vertical structural intersections.

Gold-silver mineralization is associated with a variable assemblage of pyrite-sphalerite-marcasite-pyrrhotite ± chalcopyrite ± galena ± arsenopyrite (± stibnite ± tetrahedrite ± bismuthite).  Sulphide and gangue mineralogy along with sericite alteration is reasonably characteristic of an intermediate sulphidation regime as defined by Sillitoe and Hedenquist (2003).  However, pervasive silicification is more typical of high-sulphidation deposits and minor carbonate gangue of a low-sulphidation environment.

Sulphide mineralization at Blackwater can be divided into the following types:

(i)         Disseminated - as pinhead to coarse blebby sulphide grains and aggregates typically ranging from 1% to 5% total volume of the rock, but locally exceeding this volume.  Disseminations may be uniform or irregular, with sulphides displaying an anhedral to euhedral crystal form.

Disseminations of a dark-grey, very fine-grained sulphide material called “dendritic-black-sulphide” (DBS) is common at Blackwater and may form as fine disseminations to coarse clusters, as thicker coatings to fractures, or as an irregular network of “dendritic” micro-cracks within the rock mass.

(ii)         Porosity infill - sulphides fill, rim, or replace devitrified pyroclasts, tephra, and juvenile pumiceous material.  Sulphides also commonly form parallel to compositional layering and laminations within felsic pyroclastic flows and laminated tuff units.  Mineralized amygdules and altered feldspars are also observed in the andesite flow units.

(iii)         Vein - polymetallic, anhedral to euhedral sulphide assemblages in sub-millimetre to centimetre-scale polymetallic veinlets-veins of quartz-sericite-chlorite-clay (illite) ± (iron) carbonate ± tourmaline ± vivianite.

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(iv)         Hydrothermal brecciation and related silicification - centimetre to metre-scale zones of hydrothermal brecciation, alteration, and elevated sulphide content.  These breccia zones are typically healed with silica-sericite-sulphide cement and cut by a micro-stockwork of vitric quartz ± sulphide veinlets.

(v)         Structure-related (late?) - sulphides crushed to comminuted in brittle fault breccia and gouge.

Hydrothermal alteration (and possibly contact metamorphism) has produced several superimposed alteration assemblages, including pervasive silica-sericite-clay (illite) ± biotite alteration and veinlet/fracture-controlled silica-sericite-chlorite-clay ± iron carbonate ± tourmaline.  An early (?) biotite-silica-albite±chlorite/actinolite hornfelsing event may have been significant, although mineralization in these rocks appears to be lower than in units without evident hornfelsing.  Visible native gold has been noted in some drill holes.

Secondary quartz occurs in several modes:

· Pervasive, amorphous to translucent silicification with associated illite ± sericite.  Commonly holes display intense silicification of felsic units, epiclastics, and more intermediate volcaniclastic rocks with biotite alteration of the matrix (hornfels).

· Cryptocrystalline silica replacements in felsic ash-tuff layering

· Silica cement/matrix to local hydrothermal brecciation

· Sub-millimetre vitric quartz veinlets in zones of intense silicification; commonly as a micro-stockwork.

Due to lack of outcrop, all geological interpretation has been based on drill information.  In previous resource models, interpreted geology relied heavily on indicator kriging, since it proved impossible to project perceived lithotypes between cross-sections with any degree of confidence.  The geological model discussed in Section 14 consists of two major units: andesite and fragmental volcanics.  The contact between these domains is poorly defined and often obscured by pervasive bleaching due to sericite and kaolinite alteration and silicification.  The present model is a combination of sectional interpretation and indicator kriging of the andesite domain.

Efforts are currently underway to advance the understanding and interpretation of these controls in order to develop more realistic models that can incorporate sulphide content, alteration, and silicification, as these appear to be most closely linked with gold and silver mineralization.

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Figure 7-3 is a control plan for the drill sections included as Figure 7-4, Figure 7-5 and Figure 7-6.  The sections present gold and silver grade as histogram bars representing 5 m downhole composites.  The gold composites are constrained by 0.2 g/t grade shell domains, which were generated within two main structural domains labelled 901 and 902 (see Section 14.0).  Silver composites are constrained only by the structural domains.  Lengths and average grades of some representative intervals within the gold domains are shown on the sections.

7.4 Comments on Section 7

In the opinion of the QPs, knowledge of the deposit settings, lithologies, mineralization style and setting, and structural and alteration controls on mineralization are sufficient to support Mineral Resource estimation.  The planned delineation of sulphide content, alteration, and silicification for modelling purposes will support more detailed modelling efforts.

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8.0 DEPOSIT TYPES

The Blackwater deposit is considered to be an example of an epithermal-style deposit.

Mineralization at Blackwater is hosted within felsic to intermediate composition volcanics that have undergone extensive silicification and hydrofracturing in association with pervasive stockwork veined and disseminated sulphide mineralization.

The geological setting, style of gold-silver mineralization, and associated alteration assemblage for the Blackwater deposit share the characteristics of both low and intermediate sulphidation epithermal deposit types, according to the classification system of Sillitoe and Hedenquist (2003).  Gold-silver mineralization is associated with a variable assemblage of pyrite-sphalerite-marcasite-pyrrhotite ± chalcopyrite ± galena ± arsenopyrite (± stibnite ± tetrahedrite ± bismuthite).  Sulphide and gangue mineralogy is reasonably characteristic of an intermediate sulphidation regime as defined by Sillitoe and Hedenquist (2003).  However, the massive fine-grained silicification is more typical of high-sulphidation deposits and minor carbonate gangue of a low-sulphidation environment.

A typical section showing the main features of calc-alkaline volcanic arc setting and associated epithermal and related mineralization is included as Figure 8-1.

“A calc-alkaline volcanic arc with neutral to mildly extensional stress state showing relationships between high- and intermediate-sulfidation epithermal and porphyry deposits (note that the complete spectrum need not be present everywhere).  Early magmatic volatiles are absorbed into ground water within the volcanic edifice (shown here as a stratovolcano, but it may also be a dome setting) to produce acidic fluid for lithocap generation, over and/or supra-adjacent to the causative intrusion.  Later, less acidic intermediate-sulfidation fluid gives rise to intermediate-sulfidation mineralization, both adjacent to and distal from the advanced argillic lithocap.  Where the intermediate-sulfidation fluid flows through the leached lithocap environment, it evolves to a high-sulfidation fluid to produce high-sulfidation veins or disseminated mineralization, depending on the nature of the structural and lithologic permeability.

The high-sulfidation fluid may evolve back to intermediate-sulfidation stability during late stages, supported by paragenetic relationships and lateral transitions of high- to intermediate-sulfidation mineralogy.”

Key features of these deposit styles seen at Blackwater are summarized in Table 8-1.  Figure 8-2 illustrates the hypothesized relationship of the mineralized volcanic rocks to surrounding strata at Blackwater.

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8.1 Comments on Section 8

The QPs are of the opinion that the deposit style used for exploration targeting is appropriate to the mineralization identified and the regional setting.

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9.0 EXPLORATION

9.1 Grids and Surveys

Eagle Mapping Ltd. generated detailed topography in August 2010 from an aerial survey flown on July 7 of the same year.  Topography was generated at 2 m contour intervals over an area of 5 km2 and at 5 m contours over an area of 56 km2.

Eagle Mapping performed an aerial light detection and ranging (LiDAR) survey of the Project area on August 8 and 9, 2011.  Although the area of interest (AOI) for this survey was 412 km2 in size, the survey actually covered approximately 500 km2 to buffer the true AOI for quality assurance purposes.

The LiDAR topographic data were collected using a Riegl VQ-480 laser scanner and airborne GPS/IMU.  Data was collected in 1 to 2 pulses / m2 with a +/- 0.25 m vertical accuracy and +/- 0.35 m horizontal accuracy based on ground control points.  Topographic data were provided to New Gold as contours in DWG format, ArcASCII elevation grids, and orthophotography in GeoTiff format.

9.2 Geological Mapping

New Gold has not conducted any geologic mapping or systematic surface geochemical sampling at the property.  The resource area has no significant outcrop.  As noted in Section 6.0 previously, Granges carried out 1:10,000 scale geologic mapping to the north of the deposit in 1992.  In an earlier 1984 program, Granges excavated a total of 30 hand trenches, but only one, in the northwestern part of the resource area, reached bedrock and returned an anomalous Ag grade from a grab sample.

Results from recent drilling by Richfield and New Gold indicate areas of thin overburden cover near the centre of the deposit that may support mapping using shallow trenches.

9.3 Geochemical Sampling

New Gold has not carried out any geochemical surveys.  As described in Section 6.0, previous operators performed extensive soil geochemistry testing between 1982 and 2007.

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9.4 Geophysics

During 2010, Richfield contracted Quantec Geoscience Ltd. of Toronto to conduct a Titan 24 DC resistivity and IP chargeability geophysical survey.  The objective of the study was to determine the relationship between IP chargeability and resistivity and zones of known gold mineralization within the mineral resource area to aid in geologic interpretation and drill targeting.  The survey was carried out along five 3.5 km long north-south lines spaced 400 m apart with 100 m dipole length.  In October 2011, Quantec carried out a second-phase survey, consisting of eleven 2 km north-south lines with 50 m dipole length, on behalf of New Gold.

The results of the survey indicate good correspondence between known mineralization and the Titan IP-resistivity results.  In general, zones of significant gold mineralization show a positive correlation to zones of moderate resistivity and moderate IP chargeability.  Figure 9-1 shows the survey line locations and extent.  Figure 9-2 shows the results along line 5500E, where one of the prominent chargeability features correlated well with a medium- to high-resistivity gradient, and most drill holes intersected significant gold mineralization.

9.5 Petrology, Mineralogy, and Research Studies

Selected drill samples have been sent for polished section petrographic analysis.  In 2009 and 2010, sample suites were selected for the purposes of understanding the nature of the host volcanic and volcaniclastic rocks, and the gold and silver mineralization.  Sample descriptions were performed by Vancouver Petrographics Ltd.

In 2009, Eco Tech Laboratories performed whole-rock lithogeochemical analyses with the aim of constraining the geochemical fingerprint of the host volcanic rocks by providing insight into the tectonic affinity, geochemical classification, and petrological evolution.

The Metallurgical Division of Inspectorate Laboratories completed an analysis of a drill composite from drill hole BW59.  Opaque phases identified from X-ray diffraction analysis comprised quartz, micas, orthoclase, clays and minor calcium sulphates and carbonates.  Pyrite, iron oxides (limonite, hematite, magnetite and goethite) and pyrrhotite were the main iron-bearing phases.  Mineralization identified included sphalerite, chalcopyrite, cubanite, and traces of tetrahedrite, chalcocite and dioptase.  The chalcopyrite and cubanite were observed in some samples to be tightly intergrown.  Other minerals observed included dioptase, rutile, ilmenite and traces of graphite.

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9.6 Geotechnical and Hydrological Studies

Geotechnical and hydrogeological studies completed in support of the PEA are outlined in Section 16.

9.7 Metallurgical Studies

A description of metallurgical studies is presented in Section 13.0.

9.8 Exploration Potential

The Blackwater project area offers good potential for the discovery of additional mineralization that may support mineral resource estimateion.  Work to develop this potential is ongoing and involves a combination of detailed interpretation of deposit geology from the growing body of exploration drill hole information in conjunction with expanded geologic mapping, geochemical sampling, airborne and ground-based geophysical methods, and exploration drilling, both beyond the limits of the Blackwater deposit as they are currently known, and within the greater project area.

Within the Blackwater deposit itself, potential exists to expand the mineral resource known mineralization along the western and northwestern margins.  Figure 9-3 highlights two areas where the mineralization may extend beyond the current limit of exploration drilling. In the western part of the deposit, a steep northeasterly- plunging breccia pipe, the ‘SW Breccia Pipe’, hosts higher- grade gold and silver mineralization which remains open at depth below the 1,100 meter elevation level.  Likewise in the northwestern part of the deposit, the ‘NW Silver Zone’ has been identified where potential exists to expand the known mineralization both to the north and at depth.  In this area a zone of near- surface higher- grade silver mineralization remains open to the north.  It overlies a zone of gold mineralization occurring between the 1,100 and 1,200 meter elevation levels and which remains open both to the north and at depth.  Drilling to extend and delineate these zones is planned to be undertaken as part of the future exploration program at Blackwater.

Elsewhere within the project area, surface geochemical sampling and exploration drilling have recently identified an area of favorable alteration and mineralization approximately 2 to 2.5 kilometers northwest of the Blackwater deposit.  This area remains entirely covered by post mineral glacial till, however, geochemical sampling has identified an area of anomalous gold-in-till values, which range from 50 to >10,000 ppb Au.

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Additionally, drilling in this area has intercepted moderate to strongly silica–sericite-altered volcanic, sedimentary and intrusive rocks hosting locally disseminated and stockworked pyrite-pyrrhotitie–sphalerite mineralization.  The similarity between the lithologies, styles of alteration and mineralization in this area with those observed in the Blackwater deposit warrant further exploration.

9.9 Comments on Section 9

Given the lack of bedrock exposure, no detailed surface geologic mapping has been carried out over the main deposit or surrounding area by New Gold, and geologic information has been obtained primarily by core drilling.  Areas of shallow overburden near the centre of the deposit are potential targets for bulk sampling or trench mapping/sampling programs.

Geophysical surveys have proven useful to assist in interpreting deposit geology and identifying drill targets for future exploration.

The resolution and accuracy of the surface topography as interpreted from the 2011 LiDAR survey are considered sufficient to support detailed Project studies.

In the opinion of the QPs, the exploration programs completed to date are appropriate to the style of the known mineralization within the Project area.

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10.0 DRILLING

A total of 449 core drill holes (168,276 m) have been drilled in the Project area between 2009 and May 2012 (Table 10-1).  Of this total, 134 were completed by Richfield, and 315 by New Gold.  An additional 34 RC holes were completed by Granges (refer to Table 6-1); however the precise locations of these drill holes are uncertain.  Drilling by parties other than Richfield and New Gold, referred to as legacy drilling, is summarized in Table 6-1.

A drill hole location plan showing all drilling in the Project area is included as Figure 10-1.  Figure 10-2 shows the Richfield and New Gold drilling in the Blackwater deposit area.  Figure 10-3 shows the approximate locations of the drill holes completed by other predecessor companies in the immediate area of the mineral resource estimate.

10.1 Legacy Drilling

Between 1981 and 1986, Granges drilled four diamond drill core holes for a total 311 m.  The helicopter-supported winkie drill had difficulty with the ground conditions, and some holes did not reach target depths.  Drilling focused on northern parts of the property. Results of this early drilling, including drill logs, assays, and core, have not been located.

During the 1986 season, 34 reverse circulation (RC) drill holes were completed for a total of 1,524 m.  As with data for the earlier drilling, and despite thorough search, the drill logs and assays could not be located.

During 1987, 23 diamond drill core holes (2,617.13 m) were completed on the Gold and Silver zones located over what is now recognized to be the northern portion of the mineral resource area.  Drill logs and assay data were filed for assessment and drill core is stored on the property.  Richfield re-logged the core and reviewed the data.

In 1992, 788.5 m drilled in five core holes focused mainly southwest of the Gold Zone and on the Dave claim.  Logs and assays were filed for assessment and core is stored on the property.  Richfield also re-logged this core in 2010.

Five core holes were drilled in 1994 for a total 761.68 m.  The data were filed for assessment, and the core stored.  Richfield reviewed the data and re-logged the core.

Silver Quest drilled five holes in 2005 for a total 938.65 m and another two holes in early 2006 for a total of 393.97 m.  The holes were widely dispersed.  The later drilling focused on the Quest Zone (holes DAV 05-02, DAV 06-06, and DAV 06-07).  Cumulative drilling to the end of 2006 was 5,810.79 m in 44 holes.

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Table 10-1: Drill Hole Summary Table

Series	Year	Company	Holes Drilled	Total metres	Intervals Assayed	Metres Assayed
BW0042 TO BW0059	2009	RVC	18	3,621.26	3,408	3,413.62
BW0060 TO BW0116	2010	RVC	57	21,337.03	20,034	20,219.03
BW0117 TO BW0175 (+50R)	2011	RVC	59	19,733.74	18,517	19,114.96
	Subtotal	RVC	134	44,692.03	41,959	42,747.61
BW0176 TO BW0299	2011	New Gold	125	49,153.18	46,210	46,872.92
BWMET01 TO BWMET07	2011	New Gold	7	2,281.91	1,347	1,364.73
BW0300 to BW0443.448,449	2012	New Gold	146	58,444.25	53,152	54,025.28
BW0444 to BW0478	2012	New Gold	32	11,904.42	0	0
GM12-01 to 05	2012	New Gold	5	1,800.00	1,633	1,682.04
	Subtotal	New Gold	315	123,583.76	102,342	103,944.97
	Total	All	449	168,275.79	144,301	146,692.58

Note:  No analytical data were available for 34 RC holes completed in 1986 by Granges, and their precise locations are uncertain.

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10.2 Drill Methods

Exploration drilling carried out since 2009 has been predominantly HQ diameter (63.5 mm) diamond drill core except where a reduction to NQ diameter (47.6 mm) was required to attain target depths.  The last three metallurgical holes (BWMET05 – BWMET07) were PQ diameter (85 mm) core.  Contractors and rig types used on the Project for the Richfield and New Gold drill programs are summarized in Table 10-2. Drill core is transported from drill to camp by four-wheel drive vehicle for core logging.

Table 10-2: Drill Contrator and Rig Type Summary Table

Year	Company	Drill Rig Type
2009	Falcon Drilling	F-2000
2010	Falcon Drilling	F-2000
2011	Falcon Drilling	F-2000, F-5000, F-6000
	Paycore Drilling	TITAN
2012	Falcon Drilling	F-2000, F-5000, F-6000
	Paycore Drilling	TITAN, Discovery
	Hy-Tech Drilling	S-F Tech 5000

10.3 Geological Logging

Drill core is logged in a specially-built core handling facility at the Project site.  Logging includes geotechnical, magnetic susceptibility, and specific gravity measurements taken at regular intervals.  Lithology is logged and the core prepared for systematic sampling at regular 1 m intervals.  Core sawing and sampling are the last steps in core handling.  Core is cut in half using a diamond blade rock saw, with one half of the sample interval submitted for assay and geochemical analysis and the other half returned to the core box and stored at the Project site for future reference.

Logged data are entered into LogChief™ in tables designed specifically for the Project.  Logging was carried out by Project geologists.

Magnetic susceptibility and conductivity data were measured at 10 cm increments along the core with a hand-held conductivity and magnetic susceptibility meter (GDD MPP-EM2S+Probe) and stored internally for future use.

Recovery and rock quality designation (RQD) data were measured and recorded in LogChief™.  An RQD measurement is the cumulative length of core pieces longer than 10 cm in a given core run divided by the total length of that run.  Recovery and RQD measurements were performed by company geotechnical staff.

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The lithologic nomenclature at the Project has undergone revision on two occasions since New Gold Inc. took control of the Project in June of 2011.  These steps have been taken in an effort to simplify, standardize, and clarify the nomenclature in a way that will facilitate consistency in logging, geologic interpretation, and ultimately resource modelling.  As result, the following six principal rock lithology types have been defined:  Overburden (OB), Felsic Tuff (FT), Felsic Lapilli Tuff (FLPT), Volcaniclastic (VC), Andesite (AND), and Sediments (SED).  Based on these rock types a thorough re-logging of each hole drilled prior to June 2011 has been completed and the new nomenclature applied.

10.4 Recovery

Core recovery for the 2009, 2010, and 2011 drilling programs averaged 91%, and the median core recovery was 97%.  Poor core recovery often occurs in zones of faulting and fracturing.  The average RQD value for this period was 44%.

10.5 Collar Surveys

Planned drill hole collar locations were measured in the field using hand-held global positioning system (GPS) instruments.  Locations were subsequently confirmed by Trimble differential GPS.  Elevations for the drill collars were determined by draping collar coordinates over the topography measured by the LiDAR survey.

10.6 Downhole Surveys

Down-hole surveys are performed using Reflex survey equipment, and dip angle and azimuth are recorded.  A +18.8° magnetic declination correction factor is applied to the magnetic azimuth record.  Data are entered into LogChief™ in tables designed specifically for the Project.

10.7 Geotechnical and Hydrological Drilling

Five specific geotechnical HQ holes were drilled as shown in Figure 16-4 in Section 16.0.  The total length of these geotechnical holes was 1,800 m.

10.8 Metallurgical Drilling

Seven specific metallurgical holes were drilled, four HQ holes, BWMET01–04, and three PQ holes, BWMET05–07.  The total length of these metallurgical holes was 2,282 m.  Locations of these metallurgical drill holes are shown in Figure 10-4.

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Figure 10-4: Location of Metallurgical Drill Hole

10.9 Sample Length/True Thickness

Drill hole orientations are as indicated in the example cross-sections in Section 7.0 (refer to Figures 7-4 to 7-6).  Typically drill holes intersect the tabular mineralization at a high angle, and drilled intersections approximate true widths.  Grade variations encountered in the drilling are illustrated by downhole histograms for gold and silver in the figures, and indicate areas of higher grades, low grades, and intervals of higher-grades in lower-grade zones.

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10.10 Comments on Section 10

In the opinion of the QPs, the quantity and quality of the lithological, geotechnical, collar and down-hole survey data collected in the exploration and infill drill programs are sufficient to support Mineral Resource estimation as follows:

· Core logging meets industry standards for gold-silver exploration.

· Collar surveys and down-hole surveys have been performed using industry-standard instrumentation.

· Recovery from core drill programs is acceptable to allow reliable sample data for Mineral Resource estimation.

· Gold and silver mineralization occurs within an irregularly shaped system of stockwork and disseminated sulfides that strikes approximately east-west and dips moderately to the north.  Depending on the inclination of an individual drill hole, and the local dip of mineralization, drill intercept widths are approximately equivalent to true widths.

· Drill orientations are generally appropriate for the mineralization style.

· Drill orientations are shown in the example cross-sections included in Section 7.0 and can be seen to appropriately test the mineralization.  The sections display typical drill-hole orientations for the deposits, show examples of summary assay (or mineralization intensity) values using colour ranges for assay intervals that include areas of non-mineralized and very low grade mineralization, and outline areas where higher-grade intercepts can be identified within lower-grade sections.  The sections confirm that sampling is representative of the gold-silver grades in the deposits, reflecting areas of higher and lower grades

· Drill-hole intercepts as summarized in Table 10-1 appropriately reflect the nature of the gold-silver mineralization.

· Sampling methodologies are discussed in Section 11.0, and comments on the sampling protocols are outlined in Section 11.9.

· Metallurgical recoveries are discussed in Section 13.0, and comments on the recoveries are outlined in Section 13.6.

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11.0 SAMPLE PREPARATION, ANALYSES, AND SECURITY

11.1 Sampling Methods

Drill core is transported in core boxes from the drill site to camp by four-wheel drive vehicle and logged in a specially built core handling facility, where sampling for analysis is done systematically on 1 m sample composites.  Intervals are measured and marked on the core boxes, where sample tags are stapled at the beginning of each interval.  A diamond saw is used to split the core in half lengthwise.  One half of the core is placed in a standard heavy poly sample bag identified with a re-printed sample tag in a zip-lock baggy, and the poly bag is closed with a zip tie.  The remaining half core is kept for reference in the core box and stored at the Project camp site.

Certified reference standards, blanks, and duplicates are inserted into the sample stream, and four sample bags are then placed into a larger rice bag labelled with the sample ID numbers and sealed with a numbered banker’s security tag.  Between preparation and shipment, a period of up to four days, the rice bags containing the samples are stored in a secure area behind the core cutting area.

11.2 Metallurgical Sampling

Metallurgical samples were selected from the designated metallurgical holes, and samples from numerous resource holes across the deposit as defined in early 2012.  The samples were collected and despatched from site to laboratories under the supervision of the Exploration Manager.

11.3 Density Determinations

New Gold and Richfield have collected more than 25,000 measurements of specific gravity using a water immersion method without a wax coating on drill core samples during the period from 2009 to June 2012.  New Gold selected 154 samples, from across the deposit, to be analysed at ALS to verify these field measurements.  All samples were run without a wax coating, and then a selection of 55 samples was re-analysed with a wax coating.  The results showed there is no bias between the field and laboratory methods for all but overburden samples.  Several samples identified as “overburden” returned significantly higher results from ALS, with an average density of 2.56 compared to 1.91.

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11.4 Analytical and Test Laboratories

Eco Tech, the primary analytical laboratory used by Richfield, had ISO 9001:2008 certification.  Eco Tech, located in Kamloops, BC, was acquired by ALS Minerals (ALS) during 2011.

ALS, which has been used as the primary analytical laboratory since July 2011, is ISO 9001:2008 accredited, and has ISO/IEC 17025:2005 accreditations for the following analytical techniques:

· Fire assay Au by atomic absorption (AA);

· Fire assay Au and Ag by gravimetric finish;

· Fire assay Au, Pt, and Pd by inductively coupled plasma (ICP);

· Aqua regia Ag, Cu, Pb, Zn and Mo by AA;

· Four acid Ag, Cu, Pb, Zn, Ni and Co by AA;

· Aqua regia multi-element by ICP and MS;

· Four acid multi-element by ICP and MS; and

· Peroxide fusion multi-element by ICP.

ALS analyses were completed at the North Vancouver, BC location.

Both Eco Tech and ALS are independent of New Gold.

11.5 Sample Preparation and Analysis

Standard preparation procedures are used to crush, split, and pulverize the core samples before analysis.  Protocols at the two laboratories are summarized in Table 11-1.

Table 11-1: ALS and Eco Tech Sample Preparation Protocols

	ALS	Ecotech
Crushing	90%, 2 mm	70%, 2 mm
Splitting	Rotary & Riffle, 1,000 g	Riffle, 250 g
Pulverizing	95%, 106 µm*	95%, 106 µm

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In late 2011 and into 2012, approximately 7,086 samples were prepared at ALS in Elko, NV, USA.  This amount represents approximately 5% of the total sample data set used for the mineral resource estimate in Section 14.0.  The Elko samples were pulverized to 95% passing 75 µm. Samples were also prepared at ALS laboratories in Terrace, BC, and Reno, NV, USA.  ALS Terrace and Elko have ISO 9001:2008 certification.  ALS Reno has ISO 9001:2008 certification and accreditation to ISO/IEC 17025:2005.

Assay procedures include a multi-element package (28 elements at Eco Tech, 33 elements at ALS) with aqua-regia (AR) digest (nitric and hydrochloric acids) and induction-coupled plasmaspectrometry atomic emission spectrometry (ICP AES) finish for multi-element analysis.  Both laboratories used fire assay with atomic absorption (AA) for gold, although those performed by Eco Tech were based on 30 gram pulp splits whereas those performed by ALS were based on 50 gram pulp splits.  In earlier samples, Eco Tech assayed silver by aqua regia/AA, but later switched to include silver in the overall ICP analysis.  ALS uses ICP for silver analysis.

Because of the differences in assay method and some documented discrepancies in sample identification, caution is needed in interpreting the accuracy of silver grades relative to gold grades reported in the current mineral resource estimate.

At both laboratories, overlimit ICP analyses of Ag, Cu, Pb, and Zn are conducted on analyses that returned grades of >30 g/t Ag and >1% Cu, Zn, and Pb.  For Au, overlimit fire assay (FA) analysis is conducted on analyses that return >10 g/t.

11.6 Quality Assurance and Quality Control

Quality assurance / quality control (QA/QC) protocols for the Richfield work included the insertion of certified reference material (CRM), a blank, and a quarter-hole duplicate into every set of 30 samples.  The CRMs were produced by CDN Resource Laboratories Ltd., an independent laboratory in Langley, BC.

The blanks inserted into the sample stream were originally derived from an outcrop of barren granodiorite exposed about 10 km north of the mineralized zone.  Eco Tech analyzed 13 samples of this material in 2009 before using it for sample blanks.  Limestone garden stone was later used for the blanks.

In the New Gold QA/QC programs, one CRM standard is inserted every 20 samples, and one blank is inserted every 30 samples.  As of October 2011, quarter-core field duplicates are no longer inserted.  The reference standards acquired from CDN Resources were used until the end of 2011 and were phased out of the program by February 2012.  Since then, CRMs are sourced from WCM Minerals, a division of WCM Sales Ltd. in Burnaby, B.C., and Geostats Pty, in WA, Australia.

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The reference standards were selected to match the precious metal content range and material type at Blackwater as closely as possible at the start of the exploration program.  The drill hole database was supported by some 40,000 QA/QC check assays.

11.7 Databases

The current drill hole and assay database for the Project is administered from the New Gold Vancouver office using the same Maxwell GeoServices products, LogChief™ and DataShed™, which were used during the Richfield drilling programs.

Drill hole data logged in the field are entered into a LogChief™ database specifically tailored to the Project.  LogChief™ validates the data as they are entered, and the final logs are exported and transferred to the database administrator in Vancouver for import to DataShed™, the master database.  DataShed™ performs additional data validation checks, and the administrator re-checks all logs.

The assay certificates received from both Eco Tech and ALS are delivered in a format that allows instant import to DataShed™.

Access permission for entering and editing data into the database is restricted to the New Gold Corporate Exploration Manager and the Project Database Administrator.  The database is hosted on the New Gold server, which routinely backs up every day for protection from data loss due to potential drive failures or other technical issues.

11.8 Sample Security

Samples are transported to Prince George by truck, where the driver waits with the samples in the truck until pick-up for onward shipment by a bonded courier.  Before July 2011, the Richfield samples, including the standards, blanks, and duplicates, were shipped to Eco Tech Stewart Group Laboratories (Eco Tech) in Kamloops, B.C.  Since the acquisition of Richfield by New Gold in June 2011, and the subsequent acquisition of Eco Tech by the ALS Group in July 2011, samples have been shipped to ALS Minerals (ALS) in North Vancouver, B.C.

11.9 Comments on Section 11

The QPs are of the opinion that the quality of Au and Ag analytical data collected during the Richfield and New Gold programs are sufficiently reliable (also see discussion in Section 12.0) to support Mineral Resource estimation and use of the data in a PEA, and that sample preparation, analysis, and security are generally performed in accordance with exploration best practices and industry standards as follows:

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· Data are collected following industry-standard sampling protocols

· Sample collection and handling of core were undertaken in accordance with industry-standard practices, with procedures to limit potential sample losses and sampling biases

· Sample intervals in core are at 1 m intervals; the sample intervals are considered to be adequately representative of the mineralization

· Bulk density determination procedures are consistent with industry-standard procedures, and there are sufficient bulk density determinations to support tonnage estimates

· Sample preparation for samples that support Mineral Resource estimation has followed similar procedures for the Richfield and New Gold programs, and both laboratories used the same pulverization sizing.

· Core drill programs were analyzed by independent laboratories using industry-standard methods

· Silver ICP analyses are not known with the same precision and do not have the same quality control support as gold fire assay analyses

· Typically, drill programs included insertion of blank, duplicate and CRM samples

· Data that were collected were subject to validation, using in-built program triggers that automatically checked data on upload to the database

· Verification is performed on all digitally-collected data on upload to the main database, and includes checks on surveys, collar co-ordinates, lithology data, and assay data.  The checks are appropriate, and consistent with industry standards

· Sample security has relied upon the fact that the samples were always attended or locked in the on-site sample preparation facility

· Chain-of-custody procedures consist of filling out sample submittal forms that are sent to the laboratory with sample shipments to make certain that all samples are received by the laboratory

· Current sample storage procedures and storage areas are consistent with industry standards.

Data from holes drilled between 1981 and 1994 have no documented QA/QC information, and they are not deemed acceptable for use in resource estimation.

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12.0 DATA VERIFICATION

12.1 Site Visit Verification

Mr. Simpson visited the site on December 13, 2010; September 8, 2011; November 28, 2011; and September 20, 2012.  The purpose of the visits was to review the drilling, sampling, and quality assurance/quality control procedures.  The geology and mineralization encountered in the drill holes completed to date were also reviewed.  During the site visits the Mr. Simpson verified:

· Collar locations are reasonably accurate by comparing six drill hole database collar locations with hand-held GPS readings.

- Although most drill hole collar locations have been field-surveyed by New Gold using a Trimble DGPS survey instrument, no base station exists and therefore the drill hole elevations have a low level of accuracy.

- Currently drill hole elevations are assigned by “draping” the collars onto a digital elevation model derived from LiDAR data.

· Drill hole collars are clearly marked with sturdy wooden fence posts, and the drill hole identity, orientation, and depth are inscribed onto a metal tag (Figure 10-1)

· Down-holes surveys are routinely taken at approximately 50 m intervals using a Reflex single-shot unit.

· Drill logs compare well with observed core intervals.

· Core recoveries were generally high through the mineralized zones, with some core loss noted in zones of more intense fracturing and faulting.

· Specific gravity is determined using a water immersion method where the weight of the sample in air and in water is measured with a balance beam scale.

· Assay results for four samples of drill core collected in 2010 by Mr. Simpson and submitted for assay were consistent with database values at the same intervals.

12.2 QA/QC Review

12.2.1 Standards

CRMs were used to monitor laboratory accuracy from 2009 to March 2012.

Upon receipt of the assay data, the CRM results are examined and compared with the best values for that material.  Where the results are within two standard deviations from the mean of the best values, the standard is deemed to pass.  No unacceptable standard verification results were found in 2010 or in the first half of 2011, and only one was found in 2009.  In five instances a blank was found to have been inserted into the sample stream instead of a standard, and five cases of misclassified standards being inserted were also noted.

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Before the ALS acquisition of Eco Tech, the average standard grades from Eco Tech compared closely with the best values except for CDN-GS-3E, where the average grade of 78 samples was almost 10% higher (3.26 vs. 2.97 g/t).  Examination of the CRM certificate revealed that one of the round-robin tests had returned similar levels, and so the certified value may not be completely reliable.  When performance limits were adjusted for the bias, the results were deemed acceptable with one failure.  Use of this standard was discontinued at the end of 2009.

Since switching to ALS, an increase in the standard failure rate was initially observed among some of the 2011 sample batch sequences.  The affected batch sequences were re-assayed to acceptable QA/QC performance limits.

Values from performance control charts for 19 CDN and WCM standards as tested at ALS between November 25, 2011 and February 1, 2012 were examined.  The relative biases between best value and mean value range from -3% and +4% and are not considered significant.  Similar results were achieved for the CDN standards used between 2009 and 2011.

12.2.2 Blank Samples

Upon receipt of the assay data the blanks are carefully examined and compared with accepted values.  Assay results for blanks have been uniformly below or at detection limit for gold.  For other elements the blanks show good consistency, giving confidence that the same material was being used in each instance.

12.2.3 Field Duplicates

Assay results for field duplicate pairs were compared using grade values of > 0.2 and < 10 g/t.  A total of 2,479 field duplicate pair results from 2009 to March 2012 were examined, 6% of which were duplicate half core and the rest duplicate quarter core.  Cumulative frequency plots for absolute relative difference for Au and Ag in the field duplicates were prepared.  At the 90% cumulative frequency level, the values for Eco Tech and ALS are both around 70% for Au and 60% for Ag, indicating a high level of variability between field duplicates.  A commonly-accepted value is 90% of paired duplicates with < 30% absolute relative difference.

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12.2.4 Coarse Reject Duplicates

A total of 2,066 coarse reject duplicate pair samples were assayed at Eco Tech before the ALS transition, and ALS has assayed a total of 2,389 samples since that time.  Results were evaluated by cumulative frequency absolute relative difference charts using values > 0.2 g/t Au.  The recommended value for coarse reject duplicates is less than 20% absolute relative difference at the 90% cumulative frequency limit.  The Eco Tech results yield a value of 21% at this level, representing results between 2009 and 2011.  Results for ALS show a significant drop in performance during the transition period from July to November 2011, with considerable improvement since that time.  The average for all ALS coarse reject results is 29% with statistics for December 2011 as low as 20%.

Results for Ag were similar, with absolute relative difference values around 35% at the 90% cumulative frequency limit.

12.2.5 Pulp Duplicates

A total of 409 pulp duplicate samples were assayed at Eco Tech before the ALS transition, and ALS has assayed a total of 2,304 pulps since that time.  Results were evaluated by cumulative frequency absolute relative difference charts using values > 0.2 g/t Au.  The recommended value for pulp duplicates is less than 10% absolute relative difference at the 90% cumulative frequency limit.  The Eco Tech results yield a value of 13% at this level, representing results between 2009 and 2011.  Results for ALS show a drop in performance with little change over the past eight months.  Results for Ag were similar for both laboratories, with about 30% absolute relative difference at the 90% cumulative frequency limit.

12.2.6 Check Assays

Check pairs are the results from samples assayed at different laboratories and / or using different procedures.  Check pairs are used to assess laboratory bias.

During the period that Eco Tech served as the primary laboratory, every 30th pulp reject from holes BW0060 to BW0090 was sent to Acme Analytical Laboratories Ltd., an ISO 9001 registered facility in Vancouver, B.C., for a check on the primary assays.  Acme sent the assay data to Richfield as pdf and xls files.  The statistical comparison indicates that Acme has a very minor high bias, which is not considered significant.

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12.3 Database Verification

In November 2011 and January 2012, AMEC completed data verification of the drill hole database used to support New Gold’s Blackwater mineral resource estimate.  The database verification was performed to confirm that drill hole data used to support the resource estimate are accurately transcribed.  Transcription accuracy was assessed through comparison of the drill hole database records to original collar location, down-hole survey, and assay documentation where available.  As a result of the database verification, AMEC concluded the drill hole database is reasonably free of transcription errors.  Although some errors and deficiencies were observed, AMEC concluded these would not have a material impact on the resource estimate.  The AMEC database verification work and recommendations to address observed errors or deficiencies are documented in an internal memorandum to New Gold.

12.4 Comments on Section 12

Based on the site visit observations, the QPs conclude that drilling, logging, and sampling of drill core during the Richfield and New Gold programs have been conducted in a manner appropriate to the style of mineralization present on the property.

The process of data verification performed by the QPs indicates that the data collected by Richfield and New Gold from the Project adequately reflect deposit dimensions, true widths of mineralization, and the style of the deposits, and adequately support the geological interpretations for the purpose of this Report.  The QPs are of the opinion that the analytical and database quality are adequate for the purposes of this Report.

QA/QC with respect to the results received to date for the 2009, 2010, and 2011 exploration programs is acceptable, and protocols have been well documented.

New Gold is currently investigating a possible decrease in analytical precision noted in analytical QA/QC samples analyzed by ALS since July 2011.

Drill data are typically verified prior to Mineral Resource estimation by comparing data in the Project database to data in original sources.  For most of the data, the original sources are electronic data files; therefore, the majority of the comparisons were performed using software tools.  No significant errors were found with the database that would preclude use in Mineral Resource estimation.

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13.0 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 Metallurgical Testwork

Blackwater metallurgical testing programs were conducted on samples taken from site inventory of drill core generated in the exploration drilling program (2009–2011) and from four drill holes (2011) completed specifically to generate fresh core for metallurgical testwork.

Compositing was based on oxidation state, rock type, and average Au and Ag grades.

The main objectives of the metallurgical testing programs were to characterize the mineralization styles physically and chemically and to determine the amenability of the gold mineralization to gravity separation, flotation, and cyanidation.  Results from these tests have been used as the basis for flowsheet development, process selection, and recovery forecast outlined in this Report.  The overall metallurgical program, which incorporates testwork from six laboratory locations in Canada and the USA, includes comminution, leaching, gravity, flotation, rheology, dewatering, and environmental characterization (Table 13-1).

Four broad process flowsheets were assessed for Blackwater:

· Option 1 – whole ore leach (WOL)

· Option 2 – flotation, concentrate regrind, and leach (FCL)

· Option 3 – hybrid: flotation with concentrate leach for sulphide mineralized material and WOL for oxide and transition material (Hybrid)

· Option 4 – heap leach (HL).

The PEA metallurgical test program was designed to provide information to assess the viability and advantages of these options.

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Table 13-1: Main Metallurgical Test Programs Analyzed for PEA

Program No. (Year)	Laboratory	Primary Purpose in Process Trade-off Study
RVC001 (2010)	Inspectorate Richmond, BC, Canada	Whole ore leach, gravity, flotation characterization, and mineralogy
RCV002 (2010)	G&T Kamloops, BC, Canada	Whole ore leach, gravity, flotation characterization, and mineralogy
NGD001 (2011)	G&T Kamloops, BC, Canada	Bottle roll extractions on crushed feed for heap leach characterization
NGD002 (2011)	McClelland Laboratories Sparks, NV, USA	Bottle roll extractions on crushed and ground feed for whole ore leach and heap leach characterization
NGD003 (2011)	Dawson Metallurgical Labs, Salt Lake City, UT, USA	Flotation parameter optimization and variability Mineralogy of flotation products
NGD004 (2011)	G&T Kamloops, BC, Canada	Comminution parameters
NGD005 (2012)	Pocock Industrial Salt Lake City, UT, USA	Thickening and filtration on flotation products
NGD006 (2011)	SGS Lakefield, ON, Canada	Grinding parameters
NGD007 (2011)	G&T Kamloops, BC, Canada	Gravity and whole ore leach parameter optimization and variability Flotation and regrind concentrate leaching optimization
NGD008 (2011)	SGS Vancouver, BC, Canada	Whole  ore leach parameter optimization and variability
NGD013 (2012)	G&T Kamloops, BC, Canada	Comminution parameters
NGD014 (2011)	McClelland Laboratories Sparks, NV, USA	Confirm heap leach extractions and agglomerate preparation requirements and strength characteristics

13.1.1 Crushing and Grinding Testwork

Comprehensive comminution data, which include Bond ball and rod mill work indices, abrasion indices, and JK breakage parameters, were produced during three programs: NGD004, NGD006 and NGD013.  Within these programs, 94 SMC tests have been completed to provide the SAG mill breakage parameters, along with 98 Bond ball mill work indices and 20 Bond low energy impact crushing work indices.  Data indicate that Blackwater material is hard, based on the JK Tech global database of material breakage characteristics, with oxide and transition material somewhat softer than the sulphide material.

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13.1.2 Mineralogy

As part of the RVC002 program, the mineral content of the test sample was determined using the Particle Mineralogical Analysis (PMA) function in QEMSCAN.

G&T Metallurgical found that:

· The sulphide mineral content accounted for about 4.7% of the sample weight.

· The sulphide minerals and their proportions of sulphides were:

- pyrite, 63%;

- pyrrhotite, 22%; and

- sphalerite, 11%.

· Minor amounts of copper sulphide, galena, and arsenopyrite were present in the feed sample.

Petrographic study on high gold grade composite C-2 (representing drill hole BW-59) in program RVC001 found that native gold was present primarily as finely disseminated particles (<10 µm), along with some electrum with a few larger particles of gold tellurides (calaverite-krennerite).

Additional preliminary mineralogical characterization data using QEMSCAN PMA analysis for AND, LTFF, RHY and Transition-master lithologies were obtained from the NGD008 program.  Content of free liberated gold minerals as well as its particle size distribution varied among the lithologies.  Gold minerals size distribution was found to be coarser than reported previously but was still considered fine to very fine, with top particle size reported in the range of 60 to 80 µm.  Presence of coarser gold, in the range of 200 to 300 µm, was reported only for the Transition mineralization.  Gold minerals exposure of the gold particles also exhibited differences by lithology, with the Andesite (AND lithology) sample exhibiting the highest content of free gold.

Little mineralogical information was obtained on silver.  Only insignificant traces of silver-bearing minerals, including electrum and silver-gold tellurides, were observed in RVC001.  Mineralogical observations from flotation products from the Dawson Metallurgical Laboratories NGD003 program detected the presence of silver in friebergite associated with chalcopyrite and sphalerite.  Friebergite is part of the tetrahedrite group, referenced in the literature as amongst the most difficult to leach under normal cyanide conditions, and showing partially refractory characteristics.

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13.1.3 Gravity Testwork

Gravity concentration using a centrifugal concentration device was performed ahead of flotation in program RVC002 and ahead of whole ore leach in programs NGD007 and NGD008.

Program RVC002 performed single-pass tests through a laboratory centrifugal concentrator at three nominal P80 sizes:  100, 150, and 200 µm, followed by upgrading to a gravity concentrate by panning, targeting a final mass pull of 0.1%.  The panning tailings were added to gravity tailings to make up the feed sample to flotation tests.  Only three tests included the concentrate upgrade step.

Gravity concentration procedures for NGD007 and NGD008 did not include the concentrate upgrading step.  This resulted in 2.5% mass pull average for NGD007 and 6.8% mass pull average for NGD008.  The gravity recovery associated with these large mass pulls includes not only free liberated gold but also a large portion of gold-bearing sulphides; therefore gravity recovery results from these two programs were not used in the definition of gold recovery in the PEA design criteria, as they do not reflect gravity recovery potential at industrial scale.

Concentrates produced by gravity in the NGD007 program were leached under intensive cyanidation conditions and exhibited relatively fast kinetics, reporting average extractions of 95.4% at the completion of the 72 hour test.

13.1.4 Whole Ore Leaching Testwork

With the exception of the high-grade sample tested in the RVC001-Inspectorate program, all Blackwater sulphide samples leached with relatively consistent kinetics, with an average extraction of 86.5% in 24 hours and a plateau average extraction of 87.7% at 48 hours.  Extraction results by oxidation state showed that higher extractions and faster kinetics were associated with oxide samples.  Overall results also indicated that gold leached well in the levels of oxygen provided in the standard bottle roll procedure, which includes air sparging during conditioning.

Silver extractions were consistently lower that gold extractions across all the oxidation states and lithology domains.

Average tails at 96 hour leach time were 0.11 g/t Au for sulphide and transition mineralization and 0.04 g/t Au for oxide mineralization.  Screen analysis on leach tails for program NGD002 (75 µm grind and 120 hour leach) suggests that residual gold values are finely disseminated in the residue mineral grains. Additional details provided in mineralogical characterization of WOL tails in program NGD008 confirmed that the remaining gold was very fine, under 10 µm, associated mainly with complex particles in the sulphide mineralization sample and with oxide minerals in the transition master composite sample.

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Using NGD007 data, correlations for gold and silver extraction versus grind size were weak.  On average, a 2% increase in extraction was observed as grind size was reduced from 200 µm to 75 µm.  However, the low regression co-efficient between grind size and extraction suggests that other factors, such as oxidative state, are stronger drivers of extraction in the range of conditions tested.  Similar trends were observed in the other programs.

Silver extraction exhibited the opposite behaviour, with extraction dropping with finer grinds.

The laboratory tests indicate the following trends for cyanide consumption:

· High NaCN concentration does not result in higher overall extractions or improved kinetics.

· High NaCN concentrations result in increased NaCN consumption.

The cyanide concentration optimization tests on sulphide material indicate that there is no benefit in recovery or kinetics by increasing NaCN concentration above 1 g/L in the laboratory tests.  However, higher NaCN concentration at 2.2 g/L enhanced silver extraction by about 5%.

To focus on potential gold extraction, laboratory tests were completed at a constant target cyanide concentration throughout.  In plant practice, cyanide concentrations would be staged to maintain the levels required for optimal gold extraction, and plant cyanide consumption would be expected to be lower than shown in laboratory tests.  Cyanide consumption in the plant is anticipated to be in line with industrial practice, and for the WOL flowsheet is expected to be around 1 kg/t of mineralized material.

Program NGD002 produced additional information on cyanide consumption for each oxidative state at 1 g/L NaCN concentration.  Results from this program point to a lower NaCN consumption than reported in NGD008 for equivalent gold extractions.  Based on the results of NGD002, the possibility of lower NaCN consumption is to be further investigated in the upcoming testwork programs.

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13.1.5 Flotation, Regrind, and Concentrate Leaching Testwork

Several series of testwork were completed, including flotation (with and without gravity pre-treatment), flotation concentrate regrinding, and subsequent leaching.  The work incorporated the following process variables:

· For the flotation stage: grind size, flotation time, and reagent scheme

· For the concentrate leach: regrind size, leach time, cyanide concentration, and pulp density.

Flotation

The flotation testwork was performed on a variety of samples from different lithological types and oxidation states.  Overall flotation results indicate the ability to produce a bulk flotation concentrate with recoveries averaging 90.5%, despite the variability in test conditions and resulting mass pulls.  Average mass pull to concentrate was 8.5% for NGD003 tests, 12.8% for RVC002 tests, and 15% for NGD007 tests.  In addition to test parameters and flotation time differences, variations in sulphur head grade may have contributed to the wide range of mass pulls obtained.

Analysis of the test results from RVC002 (100 to 200 µm range), NGD003 (70 to 200 µm range), and NGD007 (150 to 270 µm range) supports the following observations with regard to flotation feed grind size:

· The correlation between primary grind size and gold recovery by flotation is weak.

· A grinding range of between 100 µm and 200 µm delivers recoveries over 85%.

Overall, fine grind does not directly increase flotation recovery.  A modified reagent scheme and longer flotation time may be required for finer grind to improve recovery, as shown in the difference in recovery between NGD003 and NGD007 data.

Most of the tests completed at G&T were at grinds of approximately 200 µm.  Flotation optimization work in program NGD003 was done at a nominal grind of 100 µm.  Future phases of testwork will consider how recovery can be improved at coarser grinds.

A series of promoters was tested in combination with PAX as a primary collector and MIBC or X-133 as a frother.  CuSO4 was also used in the flotation testwork as an activator for sulphide species, specifically for sphalerite, which resulted in an increase in mass pull and a marginal increase in precious metal recovery.

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A series of tests performed at pH levels of 7, 8, and 9 indicated no significant difference in gold recoveries.  The natural pH ranged from 6.5 to 7, so future testing will be carried out at natural pH conditions.

The kinetic curves show a diminishing return trend for flotation times longer than 12 minutes, where significant increases in mass pull yield only marginal increases in gold and silver recoveries.  Corresponding trade-off analysis for laboratory flotation time showed that:

· A time of 12 minutes or less is not optimal due to lower gold recovery.

· For flotation times of 20 minutes or more, the marginal increase in gold recovery comes with a substantial increase in mass pull, resulting in increased capital costs (flotation cells and leach tanks) and operating costs (power and cyanide consumption).

· At 14 minutes flotation time, a concentrate grade of approximately 12% sulphur at 90% sulphur recovery would be expected, resulting in a mass pull of approximately 9% at average sulphur feed grades (1.3%).

RVC002 differed from the NGD003 and NGD007 programs in that it included gravity concentration ahead of flotation.  The RVC002 curve suggests that a gravity concentration stage may increase overall gold recovery, and may provide operational flexibility by buffering day-to-day variation in plant feed grade and mineralization characteristics.  RVC002 tests with gravity pre-treatment in combination with the shorter, 8 minute laboratory flotation times showed higher overall gold recoveries (93.6%) than the later tests without gravity pre-treatment (89.7% for NGD003 and 91.9% for NGD007).  Confirmation of the potential overall recovery benefit of gravity would require data generated by the Laplante E-GRG test followed by the corresponding modelling.

Concentrate Leaching

Program NGD003 included a flotation concentrate regrind series with target regrinds at P80 of 61, 35, and 18 µm.  The following were noted:

· Au and Ag kinetic curves for 35 µm and 18 µm at 2,000 ppm NaCN concentration are practically the same for the sulphide master composite sample tested.

· For Au, final extraction at 72 hours is similar for all three regrind P80 levels tested (61, 35, 18 µm).  For Ag, final extraction at 72 hours leach at 18 µm regrind shows 3% additional silver recovery compared with a 35 µm regrind.

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· Regrind sizes finer than 35 µm did not increase extractions, but would increase grinding media and power costs.  Based on these three tests, a regrind target size of approximately 35 µm appears suitable for effective leaching of gold values from flotation concentrate.

Program NGD003 included a NaCN concentration series to explore ultimate recovery at cyanide concentrations of 1, 2, 3, and 4 g/L of flotation concentrate from the sulphide material.  Average kinetic curves for concentrate from sulphide mineralization indicated that gold extraction reached a plateau between 48 and 60 hours leach time.

While the flotation concentrate from oxide material had fast leach kinetics and high extraction levels, the concentrate from transition material showed considerably slower kinetics and lower final extractions.  There was no benefit to extending the transition leaching time beyond 60 hours.  The reasons for the lower extractions for transition concentrates in this limited number of tests are not currently understood.

The effect of pulp density was also investigated, where solids by weight were varied from 20% to 45% in a 1.5 g/L cyanide solution.  Results suggest that diluted slurry moderately favours kinetics and extraction but considerably increases cyanide consumption, as expected.

The NaCN dosage series indicate that moderate NaCN concentrations of 2 g/L produces greater than 92% extraction and reasonable extraction kinetics.  NaCN consumption is expected to vary between 0.8 and 1.2 kg/t of mineralized material for nominal and peak mass pull conditions.

Combined Flotation and Concentrate Leach Performance

The metallurgical data indicate that satisfactory performance is achieved by a combination of primary grind between 150 and 200 µm, laboratory flotation time of 14 minutes, a 30 µm regrind, and 60 hour leach time.  In addition, tests indicate that gravity concentration likely contributes to gold recovery by recovering coarse and fine gold that float more slowly.  Confirmation of the overall benefit of gravity concentration will be addressed in the next stage of the Project.

Table 13-2 summarizes the average recovery by unit operation and the overall recoveries achieved in the NGD003 test program for each oxidation state.  The gold extraction on transition concentrate was significantly lower than for other oxidation states.  This was based on limited tests on transition concentrate and will be investigated further in future testwork.

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Table 13-2: Overall Gold Recovery by Oxidation State for FCL Testwork

	Flotation Recovery			Leach Extraction			Overall Recovery
Oxidation State	Au %	Ag %		Au %	Ag %		Au %	Ag %
Sulphide	91	82		93	64		85	52
Transition	84	70		85	50		72	35
Oxide	74	31		98	68		73	21

13.1.6 Heap Leach Testwork

The amenability to heap leach was tested using bottle-rolls on various crush sizes, as per industry standard.

Bottle-roll cyanide extraction tests were performed for six to seven days’ duration on crushed drill core samples in the NGD001, NGD002, and NGD014 test programs.  Tests were completed on a range of mineralization types, oxidation states, and crush sizes between 80% passing 12.5 mm and 1.4 mm.  The initial results indicate overall gold recoveries of 82% for oxide, 70% for transition, and 55% for sulphide mineralization at a 6.3 mm P80 crush size.

Subsequent column leach tests were commenced on transition and sulphide material in the NGD002 and NDG014 test programs.

The various heap leach test programs indicate the following:

· Sulphide mineralization produces gold extractions of 40% to 60% at a crush size of P80 6.3 mm.  Extraction could improve if crush size is reduced to 3.4 mm as suggested by the 7% higher extraction at this finer crush.

· Transition results exhibit higher variability, as is to be expected given the progressively weathered nature of the material from oxide to sulphide state.

· Oxide mineralization responds favourably to heap leach and appears less sensitive to crush size than other oxidation states.

· Silver extractions are likely to be overestimated due to poor precision in silver assaying for these tests.  Silver extractions are expected to be in the rage of 30% to 40%.

The high variability in the extraction data for transition material is thought to result primarily result from the high variance in sulphide sulphur content in this mineralization type and partially from laboratory precision.

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13.2 Process Alternatives Assessment

13.2.1 Flowsheet Alternatives

In January 2012, a process trade-off study was completed based on a 40,000 t/d operation incorporating the following potential flowsheets:

· Option 1 – whole ore leach (WOL)

· Option 2 – flotation, concentrate regrind, and leach (FCL)

· Option 3 – hybrid:  flotation with concentrate leach for sulphide mineralized material, and WOL for oxide and transition mineralization (Hybrid)

· Option 4 – heap leach (HL).

The WOL, FCL, and Hybrid milling options all incorporated primary crushing and coarse deposit material storage, semi-autogenous grind (SAG) and ball milling with pebble crushing, gravity gold recovery from a bleed of primary cyclone underflow stream, and intensive cyanidation of gravity concentrates.

For WOL, primary grinding cyclone overflow reports to a leach feed thickener, and thickened slurry flows to a cyanide agitation leach/carbon-in-pulp gold and silver extraction circuit.

For FCL, cyclone overflow flows to a flotation circuit, where a gold-bearing sulphide concentrate is recovered.  The concentrate is reground and thickened before treatment in a concentrate cyanide leaching and carbon-in-pulp (CIP) circuit.

The Hybrid circuit incorporates elements of both WOL and FCL to allow treatment of predominantly oxide and transition material through the WOL configuration and sulphide mineralization through the FCL configuration.

In all these milling options, the carbon from CIP is directed to an acid wash, stripping, and carbon regeneration circuit, and the recovered pregnant gold- and silver-bearing solution flows to electrowinning cells where gold and silver are electrowon and subsequently smelted into doré bullion.  CIP residue is treated in a cyanide destruction circuit using the SO2-Air process.  The heap leach option includes primary, secondary, and tertiary crushing and screening, agglomeration of crushed mineralized material with lime and cement, and transport and placement of agglomerated material onto a lined heap leach pad using overland and grasshopper conveyors, a horizontal conveyor, and a radial stacker.  The heaps will be irrigated with cyanide solution.  Pregnant solution will be collected in a pond and directed to a carbon-in-column adsorption circuit followed by carbon stripping (desorption), electrowinning, and smelting to produce a gold/silver doré product.

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An initial assessment was made of the predicted gold and silver recoveries, plus capital and operating costs, and pre-tax NPV for the listed options.  The heap leach option was found to have a pre-tax NPV of approximately $700 million less than the milling options.

The study determined that heap leach is not a viable process option for treatment of the overall Blackwater mineralized material, due mainly to:

· Low sulphide gold recovery;

· Limited ability to reduce capital cost because of the need for a fine crushing circuit, and high infrastructure costs; and

· Potential operating risks associated with a large-scale, year-round heap leach in a cold, high-snowfall climate.

The results of the heap leach column tests programs (NGD002, NGD014) indicate similar sulphide heap leach gold recoveries as assumed in the January 2012 process trade-off study, and the conclusions from the January study have not changed.  However, testwork suggests that heap leaching of smaller tonnage oxide and transition material may still be a viable process option for Blackwater.  Additional heap leach testing is underway.

Metallurgical data received since the process trade-off study release in January and corresponding analysis resulted in revised recovery assumptions for Hybrid, WOL, and FCL flowsheets.  The WOL flowsheet delivers higher LOM recoveries.

The capital and operating costs for the alternative milling options were also assessed.  Costs and recoveries available at June 2012 were input in the financial model, and it was determined that the WOL flowsheet gave the best project economics.

Considering that it has the highest pre-tax NPV and recoveries, the WOL processing option was selected for the PEA as the base flowsheet to treat 60,000 t/d at a target grind of 150 µm.  This processing option provides a more optimal balance between high metallurgical extractions and marginally higher operating costs, hence offering the best returns.

A preliminary, comparative economic optimization of the main metallurgical parameters for the WOL flowsheet was carried out to assess the pre-tax NPV impact of changes in capital cost, operating cost, and recoveries.  The results of this preliminary evaluation indicate that the optimum primary grind size for WOL is likely to be between 150 µm and 200 µm.  Leach time effect on pre-tax NPV indicates that the optimum time is of 24 hours or less, which presents an opportunity to optimize leach time in conjunction with with cyanide concentration.  Detailed metallurgical testwork is currently in progress to provide further data on grind, leach time, cyanide consumption, and metal recoveries.  The precious metals extraction estimates are based on the data collected from the NGD002 program at a P80 of 75 µm.

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13.3 Recovery Estimates

Life-of-mine recoveries from the four process scenarios assessed are summarized in Table 13-3.  The WOL recoveries are used for the PEA and are bolded.

Table 13-3: LOM Overall Recoveries by Flowsheet Option

	LOM Overall Recovery, %
	WOL	Flotation	Hybrid	Heap Leach
Au	86.8	82.6	83.6	62.9
Ag	55.2	41.8	44.9	35.9

13.3.1 Recovery Estimate for Treatment of Low-Grade Mineralized Material

The PEA mine production schedule calls for low-grade mineralized material to be stockpiled for treatment at the end of the mine life.  Limited test data are currently available for metallurgical treatment of mineralized materials with feed grades expected in this low-grade stockpile (0.3 to 0.4 g/t Au).  However, the lowest grades tested to date in sulphide do not show a decline in extraction, falling within the average.

13.3.2 High Grade Silver Mineralized Material

Sample sourcing for characterization and metallurgical testwork on these recently added resources is underway.  For the PEA, average silver recoveries by oxidation state have been assumed for mineralized material from this high-silver/low-gold zone. A nominal 20% gold recovery was assumed on the high-silver mineralized material.  Although the gold feed grade is very low, it is assumed that some of the gold is associated with the silver mineralization, allowing for a modest gold recovery from this material.  This will need to be confirmed by testwork.  The quantity of gold in the Inferred high-silver mineralized material represents a minor fraction of the gold in Indicated and Inferred Mineral Resources.  The PEA mill feed production plan has been developed for this material on the basis of a maximum extracted silver to gold ratio of 7:1 in mill product.

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13.4 Metallurgical Variability

The selection of material to define the process is adequate at this stage of reporting.  The three major oxidation states (which have dominated the design) have been covered well enough to provide information on their potential impact.  Further, the overall coverage by the sampling is adequate to provide a clear idea that the deposit material will generally perform well.

13.5 Deleterious Elements

There are no known deleterious elements that will have a significant impact on potential economic extraction

13.6 Comments on Section 13

In the opinion of the QPs, the metallurgical test work conducted to date supports the declaration of Mineral Resources based on the following:

· The metallurgical test work completed on the Project has been appropriate to establish a process route that is applicable to the mineralisation types

· Tests were performed on samples that were representative of the mineralisation for the purposes of establishing conceptual process flowsheets

· The process routes proposed all use conventional technology

· Recovery factors from the tests are appropriate to the mineralisation types, based on the available metallurgical testwork data.  On a daily basis, recovery will vary depending on feed grade and mix of mineralization types but it is expected that typically the recovery will follow the values assumed in the study; and

· Overall results indicate that the gold in Blackwater mineralization is free-milling (non-refractory).  A portion of the gold, between 17% and 32%, liberates during grinding and is amenable to gravity concentration.  In addition, a preliminary diagnostic leach test indicated that 90% to 95% of the overall gold is cyanide-soluble, with only 1% to 5% of the gold occluded in sulphides.  Silver, despite higher grades, exhibited poor amenability to low mass pull gravity concentration, and diagnostic leach tests showed moderate dissolution, in the range of 55% to 62%.  This level of dissolution is in line with observations of silver minerals such as friebergite, which has been reported to be refractory.

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14.0 MINERAL RESOURCE ESTIMATES

14.1 Key Assumptions/Basis of Estimate

The total sample database for the Project contains results from 495 core holes totalling 166,163 m drilled between 1987 and May 14, 2012.  Of these, 451 holes have been drilled since the start of 2009, including 437 holes drilled to delineate the Blackwater Mineral Resource, seven holes drilled for metallurgical testing in 2011, and five geotechnical holes completed in 2012.

Excluding the 32 holes for which results had not been received at the date of database close-out, a total of 417 holes totalling 147,619 m were used to support the grade estimations.  Data from the remaining holes were used in the density and lithologic modelling.

Due to lack of QA/QC and accurate survey information, holes drilled before 2009 were not used for statistical analysis or grade estimation of the PEA Mineral Resource, but were used in forming the lithological wire frame construction.

The PEA resource model is based on core drilling completed by Richfield and New Gold between August 1, 2009 and May 14, 2012.  The database close out date is June 8, 2012.  The date of the last drilling with assay data in the database is May 14, 2012.

Drilling is ongoing on the deposit.

14.2 Geological Models

Lithological wireframe models of the main fragmental, andesite, and sedimentary units were generated by company geologists using cross-section interpretations (Figure 14-1).  Blocks falling within the wireframes were coded to the appropriate lithology.  Re-logging of older drill holes is in process, and the development of a more comprehensive geological model is planned for late 2012.  None of the 16 drill holes that have intersected the sedimentary units at depth encountered any significant mineralization.

A bedrock surface was modelled by creating profiles based on drill hole intercepts and generating a digital elevation model.  Surfaces representing the base of the weathered oxide and oxide-sulphide transition zones were also generated and used to code the blocks (Figure 14-2).

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For preliminary statistical analysis, an exploratory data analysis (EDA) envelope was generated to exclude unmineralized or weakly mineralized drill hole data peripheral to the main deposit.  The deposit was also divided into two structural domains based on large-scale gold trends observed in previous resource estimates and contour plots of total gold accumulation in each drill hole.  These domains were numbered 901 (southwest), and 902 (northeast); and are referred to in Figure 14-3.  Domains 901 and 902 have relatively well defined orientations, with Domain 901 striking northeast-southwest with a steep dip to the northwest, and Domain 902 with a prominent lobe plunging moderately to the northwest.  Gold grade-shell domains were then generated within each structural domain by indicator kriging using a threshold value of 0.2 g/t Au.  A local, more steeply dipping orientation of the mineralization is present in the northern portion of the predominantly flat dipping Domain 902.  A higher-grade Ag domain (GS10) lies along the north side of the deposit close to surface.  An indicator-kriged grade shell using a threshold value of 10 g/t Ag was generated as a domain.

14.3 Exploratory Data Analysis

Within the EDA envelope, Au and Ag show a weak positive correlation (correlation coefficient = 0.28), and a linear regression yields a very low R2 value of -0.03.  As a result, geostatistics and block grade estimation for Ag were not constrained by the 0.2 g/t Au grade shells but rather by the outer structural domain boundaries (with the exception of one higher-grade Ag zone lying on the north side of the deposit close to surface).

Cumulative frequency distributions for the Au samples within the 0.2 g/t Au grade shell domains used to constrain block model grade estimation were assessed.  Sample populations are highly skewed approaching lognormal distribution with no significant bimodality evident.  There is evidence of a minor, higher-grade Au grade population as indicated by a change in slope in the cumulative probability curve above a level of 2 g/t Au.  The coefficient of variation (CV) for Au is 3.6, which is considered quite high.

It is recommended that future resource modelling work evaluate the potential benefits of reducing the CV and further constraining the influence of higher-grade gold samples via alternative estimation methods such as indicator kriging.

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14.4 Density Assignment

A total of 18,792 density measurements made on core sampled between 2009 and June 8, 2011 have been used for density modelling.  The samples were statistically analyzed by weathered zone, and outliers were removed.  The resulting data were used to interpolate density values using the inverse distance squared method (ID2).

Separate interpolations were carried out within the oxide, transition, and sulphide zones using hard boundaries.  Soft boundaries were used between the andesite and fragmental lithologic domains.  An isotropic search ellipsoid was used with a maximum search distance of 150 m.  A minimum of two and maximum of nine samples were required to estimate a block.  A maximum of three samples was permitted from a single drill hole.

Blocks that were not estimated were assigned the mean value for the corresponding lithology and material type (Table 14-1).  The density of the overburden was assigned an assumed value of 2.0 g/cm3.

14.5 Grade Capping/Outlier Restrictions

Before compositing sample assays, grade distribution in the raw sample data was examined to determine if grade capping or special treatment of high outliers was warranted.  Cumulative log probability plots were examined for outlier populations, and decile analyses were performed for Au within the zone domains.

For the 0.2 g/t Au grade shell domains, the last decile for Au contained between 57% and 58% of the metal content, and the top centile contained between 26% and 27%.  After reviewing the probability distribution, it was decided to cap Au grades at a level of 40 g/t.  The capping affects a total of 62 samples and removes an expected 4.3% of the metal.

For Ag outside of the 10 g/t Ag grade shell domain, the last decile contained 54% of the metal, and the upper centile contained 19% for domains 901 and 902.  Within the grade shell domain, the last decile contained 32% of the contained metal, and the upper centile contained 5.7%.  A cap grade of 150 g/t was selected for Ag within the structural domains but outside of the 10 g/t Ag grade shell domain.  This affects 74 samples and removes and expected 2.9% of the metal.  Silver samples within the 10 g/t Ag grade shell domain were not capped.

The resulting caps are summarized in Table 14-2.

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Table 14-1: Density Assignments by Lithology

Code	Lithology		Material Type
		All	Sulphide	Transition	Oxide
1	Overburden	2.00	-	-	-
10	Andesitic Volcanics	2.76	2.76	2.73	2.68
20	Felsic Volcanics	2.71	2.73	2.66	2.60
30	Sediments	2.71	2.71	-	-

Table 14-2: Capping Statistics by Domain - Au

Domain	Count	Uncapped Mean (g/t)	Capped Mean (g/t)	% Difference
901	35635	0.879	0.848	-3.5%
902	24655	0.810	0.764	-5.7%
ALL	60290	0.851	0.814	4.3%

Block grades were estimated using uncapped gold grades to determine the amount of metal removed by the grade capping.  The overall gold metal loss amounted to 3.5%, which is below that predicted by the capping study but still considered reasonable for the style of mineralization.  The difference between the predicted and model average grades is attributed to spatial variability.

A comparison of blocks using capped versus uncapped Ag grades gave an average capped grade of 3.5 g/t compared to an uncapped average of 3.7 g/t, or a 5.4% metal loss.  This is greater than the 2.9% difference predicted by the capping study but only reflects a drop of 0.1 g/t in the average grade differential.

14.6 Composites

Down-hole composites for Au were created within the domains using the “best fit” method.  This procedure produces samples of variable length but of equal length within a contiguous drill hole zone, ensuring the composite length is as close as possible to the nominated composite length.  In this case, the nominated length was set at 10 m. Samples within domains 901 and 902 were capped prior to compositing at 40 g/t Au.

Fixed-length down-hole sample assay composites of Ag were created using a 10 m interval within the broader structural domains.  Samples were separated into two populations, those within a 10 g/t grade-shell domain and those outside of the domain.  Composites that fell outside of the higher-grade domain were capped at 150 g/t Ag.

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The combination of capping and compositing reduced the CV for Au from 3.93 in the raw sample data to 1.59 in the capped and composited dataset.  Although this is still considered quite high, it is not unusual for precious metal deposits and is a considerable improvement over the CV of the initial sample intervals.

The combination of capping and compositing reduced the CV for Ag from 3.6 to 1.45 for the population outside the 10 g/t Ag grade shell domain.  Within the higher-grade domain the CV dropped from 0.9 to 0.7.

The interpreted contact between the fragmental unit and the andesite was partly based on mineralization, and grade estimates were confined to the fragmental unit domain.  Based on drilling to the database close-out date, the modelled andesite wireframe did not contain mineralization that was considered to have sufficient grade to warrant estimation.

Secondary elements associated with Au-Ag mineralization, including As, Ca, Cd, Cu, Ni, Pb, Total S, Sb, and Zn, were estimated by the inverse distance method to the second power (ID2) in a single domain using 10 m down-hole composites.  These elements were required to support development of an acid rock drainage (ARD) model.

14.7 Variography

Indicator variograms within each of the two structural domains were generated using an indicator threshold of 0.2 g/t Au.  These were modelled to determine the anisotropy of the structural domains.

To determine kriging parameters, search parameters, and anisotropy, directional pairwise relative semi-variograms for Au and Ag were modelled using composites falling within the corresponding domain constraints.  In the case of Au, composites were restricted to the two indicator grade shells.  For Ag, the composites were combined into a single population, as the number of sample pairs within the 10 g/t Ag grade shell was inadequate to model variograms.

14.8 Estimation/Interpolation Methods

A block model was created in Gemcom-Surpac Vision© software using a block size with dimensions of 10 m x 10 m x 10 m.  The model blocks were first coded by the partial percent within the zone domains and below topography.  Lithologic codes, material types, and density values were then assigned.

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Indicator kriging was carried out using Leapfrog3d© software to generate grade shell solid models within each of the two structural domains at a threshold of 0.2 g/t Au.

Gold and Ag grades within the corresponding zone domains were estimated in three passes using ordinary kriging and the inverse distance squared weighting method (ID2).  A single pass nearest-neighbour estimate was also carried out for use in model validation.  Search parameters are outlined in Table 14-3 and Table 14-4 for Au and Ag respectively.  The anisotropy conforms to the search ellipsoids derived from the variogram models.

Table 14-3: Au Grade Model Search Parameters

Domain	Pass	Search Distances				Composites Used			Max per Hole	Topcut g/t Au
		Major	Semi-Major	Minor		Min	Max
901	1	36	36	19		3	9		2	40
	2	75	75	41		3	9		2	40
	3	150	150	81		3	9		-	40
902	1	36	36	24		3	9		2	40
	2	75	75	50		3	9		2	40
	3	150	150	100		3	9		-	40

Table 14-4: Ag Grade Model Search Parameters

Domain	Pass	Search Distances				Composites Used			Max per Hole	Topcut g/t Au
		Major	Semi-Major	Minor		Min	Max
Outside GS10	1	36	36	28		3	9		2	150
	2	75	75	58		3	9		2	150
	3	150	150	115		3	9		-	150
GS10	1	36	36	28		3	9
	2	75	75	58		3	9		2	-
	3	150	150	115		3	9		-	-

Note:  GS10 = 10 g/t Ag grade shell

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14.9 Block Model Validation

A comparison between the grade shells and nearest-neighbour (NN) model of the same grade indicator showed that the grade shell volume is very similar to the NN model volume and is therefore considered to be unbiased.

Model verification was initially carried out by visually comparing blocks and sample grades in plan and section views.  The estimated block grades showed reasonable correlation with adjacent composite grades.

The amount of grade smoothing in the block model grade estimates was evaluated using change-of-support from composites to blocks.  A Hermitian Polynomial change-of-support method (Herco) was used to compare the distribution of the Herco-corrected declustered composites (NN block grades) with the block grades for the estimated models by plotting grade-tonnage curves.  In general, the estimated models should be slightly higher in tonnage and slightly lower in grade when compared to the Herco distribution at the selected cut-off grade.  The ID2 model closely matches the Herco-corrected grade-tonnage curve based on a block size of 20 m x 20 m x 20 m.  The current 10 m x 10 m x 10 m block model has the variability (and therefore internal dilution or selectivity) appropriate for a 20 m x 20 m x 10 m selective mining unit (SMU) at cut-off grades between 0.2 g/t and 0.3 g/t Au.

Block grades were estimated by kriging, inverse distance (ID2), and the NN methods.  A comparison of global mean values within the Au grade shell domains shows a reasonably close relationship with samples, composites and block model values.

Swath plots were generated to assess the model for global bias by comparing kriged, ID2, and NN estimates on panels throughout the deposit.  Results show a reasonable comparison between the methods, particularly in the main portions of the deposit.

14.10 Classification of Mineral Resources

Consideration was given to classifying blocks estimated in the first resource estimation pass as Measured.  However, the structural and geologic modelling is still at an early stage, and some assay/data quality issues are still under investigation.  Consequently, it was decided to delay classifying any blocks as Measured at this stage.

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Blocks were initially classified as Indicated if the block centroid was within 38.5 m of the closest hole, the second-closest hole was within 50 m, and the average distance from the two closest holes was also less than 38.5 m.  This corresponds to a drill spacing of 50 m x 50 m with a 10% contingency buffer.  The Indicated classification was then smoothed to eliminate small areas of Inferred surrounded by Indicated blocks.  Finally, Indicated blocks adjacent to the outer grade shell boundaries were downgraded to Inferred to reflect the uncertainty of grade estimates along the hard boundaries.

All other estimated blocks were classified as Inferred.

14.11 Reasonable Prospects of Economic Extraction

To assess reasonable prospects for eventual economic extraction a Lerchs-Grossmann (LG) optimized pit was prepared using general economic and technical assumptions listed in Table 14-5, was used to constrain estimated blocks.

For deposits like Blackwater, which are amenable to open pit mining methods, leading practice is to declare resources within constraining pit shells that are often generated using higher prices than those used for designing pits for which Mineral Reserves are declared.  These pits also are typically generated ascribing positive value to Measured, Indicated, and Inferred blocks that defray mining, processing, and general and administrative costs.  Furthermore, such conceptual pits should have net values (operating profit) that will return initial and sustaining capital investment.  The economic assumptions for the LG optimized mineral resource pit outlined in Table 14-5 are slightly more conservative than the economic assumptions used in the mining schedule in Section 16.0.

Table 14-5: Lerchs-Grossmann Optimized Pit Economic Assumptions

	Parameter
Pit Slope	40o
Mineralized Material Mining Cost	$1.52 / tonne
Processing Cost	$7.11 / tonne
G&A Cost	$0.95 / tonne
Waste Mining Cost	$1.90 / tonne
Gold Recovery	83.6%
Silver Recovery	44.9%
Gold Price	$1300/oz
Silver Price	$24/oz
Exchange Rate USD:CAD	1.00

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14.11.1 Determination of Cut-off Grade

Using the assumptions listed in Table 14-5, a 0.23 g/t gold-equivalent cut-off would return $8.06/t for Blackwater open pit mineralization.  This covers the assumed processing and general and administrative costs (i.e., these figures represent the marginal cut-off).  The marginal cut-off is based on the generally accepted practice that a decision is made at the pit rim if mined material above the marginal cut-off grade will lose less money if it is sent to the mill rather than if it is sent to the waste dump.  It is considered to be potentially economic if it contains a value that is greater than the costs to process it.

14.11.2 Gold Equivalency

A gold equivalent (AuEq) grade was calculated using assumed metal prices of US$1,300/oz and 83.6% metallurgical recovery for gold and US$24/oz and 44.9% metallurgical recovery for silver.  The gold equivalence formula is as follows:

AuEq g/t = Au g/t + (Ag g/t)(Revenue Ratio)
Revenue Ratio	= Gross silver revenue per tonne/gross gold revenue per tonne
	= US$ 24/oz Ag x 44.9% / US$ 1,300/oz x 83.6% = 0.0099

A gold equivalent was not used in the preparation of the gold grade shell used to constrain grade estimation.

14.12 Mineral Resource Statement

Mineral resources have been classified using the definitions set out in CIM (2010).  Mineral resources have an effective date of July 27, 2012.  The Qualified Person is Ronald G. Simpson P.Geo.  Mineral resources are summarized in Table 14-6, which presents the basecase estimate at an AuEq cut-off grade of 0.3 g/t AuEq, and sensitivity estimates around this basecase to show the sensitivity of the mineral resource estimate to changes in AuEq grades.

The distribution of resource tonnes by material type for the base case is 3% within the oxide zone, 9% within the transition zone, and 88% in the sulphide zone.

These data differ slightly from those reported in the New Gold July 18, 2012, News Release because the LG resource pit shell and gold equivalent cut-off were based on an assumed silver price of $23/oz, which has been increased to $24/oz for this resource statement.  This silver price change marginally changes the gold equivalent formula.

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Table 14-6: Blackwater Project Indicated Mineral Resource Summary (Effective Date July 27, 2012, Ronald G. Simpson P.Geo)

Cut-off Grade	Tonnes 000's	Grade			Contained Metal
AuEq g/t		Au g/t	Ag g/t	AuEq g/t	Au M oz	Ag M oz	AuEq M oz
0.20	285,241	0.836	4.2	0.878	7.67	38.52	8.05
0.25	280,379	0.847	4.2	0.889	7.64	37.86	8.01
0.30	267,145	0.876	4.3	0.919	7.52	36.93	7.89
0.35	249,371	0.917	4.5	0.962	7.35	36.08	7.71
0.40	230,572	0.963	4.6	1.009	7.14	34.10	7.48
0.45	210,646	1.017	4.7	1.064	6.89	31.83	7.21
0.50	191,572	1.074	4.9	1.123	6.61	30.18	6.92

Table 14-7: Blackwater Project Inferred Mineral Resource Summary (Effective Date July 27, 2012, Ronald G. Simpson P.Geo)

Cut-off Grade	Tonnes 000's	Grade			Contained Metal
AuEq g/t		Au g/t	Ag g/t	AuEq g/t	Au M oz	Ag M oz	AuEq M oz
0.20	133,219	0.642	7.0	0.711	2.75	29.98	3.05
0.25	128,572	0.659	7.0	0.728	2.72	28.94	3.01
0.30	120,478	0.687	7.3	0.759	2.66	28.28	2.94
0.35	110,383	0.724	7.5	0.798	2.57	26.62	2.83
0.40	98,937	0.771	7.8	0.848	2.45	24.81	2.70
0.45	88,061	0.822	7.9	0.900	2.33	22.37	2.55
0.50	77,525	0.882	7.7	0.958	2.20	19.19	2.39

Notes to accompany Mineral Resource tables:

1.	Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability
2.	Mineral resources are amenable to open pit mining methods and have been constrained using a Lerches-Grossmann optimized pit.
3.	Assumptions include US$1,300/oz Au, US$24/oz Ag, 83.6% Au recovery, 44.9% Ag recovery, $1.52/tonne mining cost, $1.90/tonne waste mining cost, $10.52/tonne process and G&A cost.  No allowances have been made for mining losses and dilution. Pit slope angle of 40°.
4.	The base case gold equivalent (AuEq) cut-off (bolded) is greater than the conceptual marginal cut-off of 0.23 g/t.
5.	AuEq= US$ 24/oz Ag x 44.9% / US$ 1,300/oz x 83.6%
6.	Gold analyses are performed by fire assay/AA finish methods and silver analyses are performed by Induction Coupled Plasmaspectrometry (ICP). Silver ICP analyses are not known with the same precision and do not have the same quality control support as gold fire assay analyses
7.	Totals may not sum due to rounding as required by reporting guidelines

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14.13 Factors That May Affect the Mineral Resource Estimate

Areas of uncertainty that may materially impact the Mineral Resource Estimate include:

· Commodity price assumptions;

· Assumptions that all required permits will be forthcoming;

· Pit slope angles;

· Metal recovery assumptions; and

· Mining and process cost assumptions.

Relogging of older drill holes may result in local changes to the lithological interpretations, on which the current block model, and resource estimate, is based.

To support a declaration of higher confidence Mineral Resource categories, the following will need to be addressed:

· Additional quality control support for silver results;

· Silver metallurgical recovery results for high Ag-low Au zone modelled as domain GS-10; and

· Additional infill drilling.

· There are no other known factors or issues that materially affect the estimate other than normal risks faced by mining projects in the province in terms of environmental, permitting, taxation, socio-economic, marketing, and political factors.  New Gold is not aware of any known legal or title issues that would materially affect the Mineral Resource estimate.

14.14 Drilling Completed Since Database Closeout Date

Drilling on the Project is ongoing.  From the PEA resource cut-off up to September 14, 2012, 377 additional holes and 100,363 m of drilling has been completed.  Complete or partial analytical results for an additional 76 holes (27,968 m) were received between the closeout date and August 28, 2012.  Of these, 69 holes (24,892 m) were infill and 7 (3,076 m) were step-out or boundary definition holes.  A review of the drill data available as at August 28, 2012 was performed to compare the results with the block model.  Although incorporation of these drill holes in an updated block model will result in local differences in interpretation of the geological and grade shells, visual inspection of the drill holes against the block model indicated the results were generally consistent with the current interpretations as used in the model.

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A review of drill log information from post database cut-off holes for which assays had not been received was also carried out.  No obvious geological features were identified that would significantly change the interpretations used to construct the block model that supports the mineral resource estimate.

14.15 Comments on Section 14

The QPs are of the opinion that the Mineral Resources for the Project, which have been estimated using core drill data, have been performed to industry best practices (CIM, 2003), and conform to the requirements of the 2010 CIM Definition Standards.

· The model domains are considered reasonable for this stage of the Project and are being improved as more drill data are obtained.

· There is sufficient density and quality of sample data to support the present resource classifications.

· The amounts of metal removed by grade capping are considered appropriate for the style of mineralization, and a check of metal removed in grade estimation is consistent with composite capping.

· Future resource models may be improved through the application of alternative estimation methods such as restrictive or indicator kriging in order to further reduce the CV and constrain the influence of higher-grade gold samples.

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15.0 MINERAL RESERVE ESTIMATES

This section is not relevant to this Report.

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16.0 MINING METHODS

The mining section is preliminary in nature and partly based on Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the PEA based on these Mineral Resources will be realized.  Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.  Dates in this section are for illustrative purposes only, as a decision to proceed with mine construction will require formal approvals from appropriate regulatory authorities and from New Gold’s Board, and and will require support from additional, more detailed, studies, and declaration of Mineral Reserves.

16.1 Pit Design

16.1.1 Mine Plan

Conventional open pit mining will be employed to mine the Blackwater deposit.  The LOM production will be 255.9 million tonnes (Mt) from the Indicated category and 100.0 Mt from the Inferred category.  Mill feed will be delivered to a primary crusher that feeds a whole ore leach (WOL) gold–silver recovery plant.  The subset of the Mineral Resources used in the mine plan is summarized in Table 16-1.

Table 16-1: Subset of Mineral Resources Used in Proposed Production Schedule

Category	Mt	Au (g/t)	Ag (g/t)	In-Situ Au oz (Million)	In-Situ Ag oz (Million)
Indicated	255.9	0.87	4.3	7.17	35.3
Inferred	100.0	0.65	7.9	2.08	25.5

The milling rate is 60,000 metric tonnes (tonnes) per day (t/d), or 21.9 Mt per year (Mt/a).  The mine life is 16.4 years after commercial production begins.  The total tonnage mined from the open pit is 1,195 Mt at an initial daily mining rate of 169,000 t/d and increasing to 273,000 t/d during Years 3 through 10 before declining after the peak waste mining period.  A total of 7.2 million ounces of gold and 35.3 million ounces of silver are estimated to be contained in the Indicated category of mill feed tonnage during the life of the mine.  An additional 2.1 million ounces of gold and 25.5 million ounces of silver are estimated to be contained in the Inferred category of mill feed.

The mine plan includes both Indicated and Inferred Mineral Resources to make up the mill feed tonnage.  Material not processed is waste.  Some of the mill feed, referred to a low-grade mineralized material, will be stockpiled in the early years for re-handle into the mill at the end of the mine life.  Another sub-set of the mill feed is within a high silver–low gold area in the northern part of the deposit, referred to as high silver mill feed.

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For the PEA mine plan, no dilution or mining loss has been applied to the tonnage and grades reported from the Mineral Resource block model.  Table 16-2 summarizes the attributes of the model variables pertinent to the mine plan.

Table 16-2: Model Variables used in PEA Mine Plan

Model Variable	Description
Gold	Block Gold grade in g/t
Silver	Block Silver grade in g/t
Class	Confidence code: 1=measured (none assigned), 2=indicated, 3=inferred
Lith	Lithology code: 1 = overburden, 10 = andesite, 20=Lapilli Tuff, 30 = silica flooded material, 50 = rhyolite
Sg	Specific gravity (used to assign tonnage per block x 1,000
otype2	Mill feed types based on oxidization: 1 = sulphide, 2 = transition, 3 = oxide, 4 = overburden (based on lith code), 5 = north high silver zone (otype2 variable adds  north high silver zone  to the original otype variable in the resource block model)
Claim	Claim owner (assigned to nearest whole model block): 1 = Joint Venture (north) 0% royalty; 2 = Dave (southwest) 1.5% royalty, 3 = Jarrit (southeast) 1.0% royalty
ard_type	Acid rock drainage type: 1 = PAG1, 2 = PAG2, 3 = NAG3, 4 = NAG4, 5 = overburden (from lith code), 0 = unassigned
hi_silver	Blocks identified as high silver, low gold zone in north model area (benches 1510 to 1380), 1 = inside high silver zone
recAu6wol	Recovered gold grade using July 2012 recoveries by otype2, zero recovery if au < 0.07 g/t
recAg5wol	Recovered silver grade using July 2012 recoveries by otype2
nsrWOL6rz	Net Smelter Return (NSR) based on US$1,200/oz gold, US$20.00/oz silver, July 2012 recoveries, July 2012 royalties, blocks outside of US$1,300/oz mineral resource pit shell set to zero NSR

16.1.2 Consideration of Net Smelter Return Values

A NSR value is calculated for each block in the model with an assigned gold grade.  The NSR value is used to express the value of a block because both gold and silver contribute to the value and their recoveries vary by the mill feed type (based on oxidation state), and because different royalties apply to the various claim blocks.  The NSR value therefore incorporates the gold and silver grades, the unique metal recoveries in the mill, metal prices, the various royalties, and post-property doré costs.  

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The NSR value is expressed as $/tonne ($/t).  Table 16-3 summarizes the inputs to the NSR calculation.

Table 16-3: NSR Inputs

Input	Description
Metal grades	Gold and silver block model grades in g/t
Metal prices and Exchange rate	Gold, US$1,200/oz, Silver, US$20.00/oz, US$1.00:CAD$1:00
Metal recoveries	Gold: oxide = 91.9%, transition = 86.7%, sulphide = 86.7%, gold < 0.07 g/t = zero recovery Silver: oxide = 53.8%, transition = 59.6%, sulphide = 51.9%
Dore charges	1.5% of NSR
Royalties*	(Davidson) = 0.00%, Dave = 1.50%, Jarrit = 1.00%

The high-silver zone in the northern part of the deposit has been identified as a unique mill feed because of the high silver grades (average 45.5 g/t in mill feed) and low gold grades (average 0.067 g/t).  The recovery of gold at these low grades has not been tested, thus for the PEA mine schedule, the gold recovery is set to zero for gold values less than 0.07 g/t.  However, in the economic analysis, a 20% recovery has been assigned to the material.

The NSR considerations are based on the assumption that the royalty on the Davidson Claims has been bought out.

The metal prices used are expressed in US dollars.  For mine planning purposes, the conversion between US dollars and Canadian dollars is 1:1, thus all NSR $/t values are equally expressed in US or Canadian dollars.

16.1.3 Pit Slope Angles

Knight Piésold (KP) has provided design criteria for overall and inter-ramp slopes for the pit walls (refer to Section 16.7.1).  The inter-ramp slopes are 27° in the overburden and range from 38.5° to 46°, in rock, depending on the pit wall orientation.  The overall slope angles (including flattening for the overburden) range from 34° to 43°, depending on the pit design sector.  The inter-ramp and overall slope angles by pit design sector have been incorporated into the pit definition floating cone runs and in the mining phase and final pit designs.

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16.1.4 Pit and Phase Design

The final pit and internal mining phases are based on floating cone geometries developed from several iterations of combinations of costs, recovery models, and metal prices.  The NSR value is based on the metal prices, metal recoveries, and royalty agreements with the three royalty holders over the property.  The sum of process, G&A, and TSF operating costs ($9.19/t, rounded to $9.20/t) is used as the “internal cut-off” for the tabulation of material to be delivered to the mill (refer to Table 16-4).  All tonnage below the internal cut-off is considered waste.

Table 16-4: Input Costs for Pit Definition

Cost Centre	$/t milled
Process	7.74
Property G&A	0.95
TSF	0.50
Sub-total	9.19
Mining:	$/t mined
Base cost for mill feed at 1480 bench, haul to crusher	1.64
PAG1 haul to TSF for sub-aqueous storage	1.98
PAG2 haul to TSF for sub-aqueous storage or Dam C embankment	1.99
NAG3 average of haul to TSF embankment or near pit dump	1.87
NAG4 average of haul to TSF embankment or near pit dump	2.00
Overburden average of haul to TSF embankment or near pit dump	1.54
Lift cost per bench per tonne below 1480 elevation	0.027

Note:  Slope angles are variable and range from 34º to 43º, overall.  Internal discount rate is 0.5% per 10 m bench

16.1.5 Pit Definition Floating Cone

A suite of cone runs based on percent of NSR was generated to evaluate the development of the pit from a cone based on 40% of NSR value to 100% of NSR value.  The net value contained within the cone geometry continues to increase up to 90% of NSR, where it flattens off with little increase in value per tonne as the cone shell increases in size.  This is in part because of the increase in waste tonnage per tonne of mill feed.  The 90% of NSR cone geometry was selected as the guide for the final pit design geometry.

16.1.6 Phase Design

The final pit is sub-divided into four mining phases based on the progression of the floating cone geometries as the percent of NSR value is increased.  A lesser number of phases were investigated but these were found to increase tonnage movement early in the mine plan and hence appeared less favourable.  The PEA phase design sequence has three phases on the south wall which smoothed out the mining rate and extends the peak mining years from four to eight years, thus better utilizing the equipment fleet.

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The pit and phase design parameters include:

· Knight Piésold recommended inter-ramp slopes by design sector and the requirement that no pit wall be more than 200 m high without a ramp or catch bench.

· Haulage ramps are a maximum of 10% in grade and 33 m wide including allowances for berms and ditches.

· The mining width between phases on a bench ranges from a minimum of 110 to 150 m up to more than 200 m on the south wall, with some short sections at 75 m.  The north side between the Phase 2 and Phase 3 pit walls is up to 700 m in width north-south.

During the phase and final pit design process, haul roads are added and pit walls are smoothed compared to the floating cone geometry.  The final pit design includes 98.4% of the cone mill feed tonnage and 3.3% additional waste tonnage.

Phase 1 is located in the central pit area and mines the shallower high-grade mill feed with the pit bottom at the 1430 elevation.  Phase 2 expands Phase 1 in all directions (except to the northeast), and the pit bottom is lowered to the 1300 elevation.  Phase 3 expands the pit to the north, picking up the north high-silver zone and a deep high-grade zone with the pit bottom of Phase 3 at the 1170 elevation.  Phase 3 is the final pit limit in the north-central part of the pit.  Phase 4 extends the pit out to the final pit limits on the east, south, and west walls, and the pit bottom is at the 1130 elevation.  The final pit configuration is shown in Figure 16-1.

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16.1.7 Pit and Dump Maps

Annual pit and dump plans have been developed to show the mining and dump advances.  These are used to measure the haulage profiles for the various material types from sources in the pit to destinations.

16.2 Proposed Mine Production Schedule

The mine production schedule is developed to deliver 60,000 t/d of mill feed to the primary crusher, provide construction materials to the TSF embankment, and maintain mining flexibility to ensure consistent mill feed.  The total material mined per day peaks at 273,000 during Years 3 through 10 assuming on average 355 operating days per year (10 days lost due to weather or holidays).

Inputs to the production schedule include:

· Maximize NSR value and gold head grades in the early years by using an elevated cut-off strategy.

· Stockpile lower-grade material when needed to maintain higher gold head grades to the mill without developing an unrealistic mine plan in terms of total tonnage movements and bench descent rates within the mining phases.

· Stockpile the high silver–low gold tonnage from mining Phase 3 of the open pit so as not to dilute the gold head grade or exceed a 6 to 1 silver to gold ounce ratio by production.

· Keep the number of benches mined in a phase to eight or less in any one year.

· Provide appropriate waste material types for construction activities at the TSF.

The mine production schedule is the culmination of multiple schedule trials at various mill throughput rates ranging from 35,000 t/d to 70,000 t/d.  Several schedules were also completed to test sequences of cut-off grades and the impact on total material movements and the amount of tonnage that would have to be stockpiled and re-handled to the primary crusher.

The NSR cut-off grade is elevated during preproduction through Year 7 and ranges between $11.00 to $17.50/t NSR.  Starting in Year 8, the NSR cut-off grade is at $9.20/t, which is the internal cut-off grade.  The cut-off strategy is shown in Figure 16-2.  The tonnage between $10.00 and the mill cut-off grade during preproduction through Year 7 is stockpiled (26,873 kt represented in green in Figure 16-2) for processing after mining from the pit is completed.  Stockpile material could also be used as a mill feed if pit production were interrupted.  The $10.00/t cut-off was selected to account for re-handle costs from the stockpile to the crusher.

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Figure 16-2: Summary LOM NSR Cut-off Grade Profile

The proposed PEA mine schedule is shown in Table 16-5.

The planned mine schedule starts with 30 million tonnes (30,000 kt) mined during a preproduction period.  Of this, 4,234 kt of Indicated and 1,549 kt of Inferred mill feed is stockpiled for processing in Year 1 and 701 kt of Indicated and 501 kt of Inferred low-grade mill feed (NSR value greater than $10.00 and less than $15.00/t) is stockpiled.  The 23,015 kt of waste is partly used for TSF construction.  During Year 1, 19,710 kt of mill feed is processed (90% of mill capacity to allow for a ramp-up period), including the mill feed stockpiled during the preproduction period.

The high silver–low gold mill feed material located in the north side of the Blackwater deposit and mined as part of Phase 3 is stockpiled during Years 3 through 7.  This tonnage amounts to 9,975 kt and is processed during Years 8 through 14 at a rate of 1,500 kt/a.  Stockpiling of this material reduces the impact of the 0.069 g/t average gold head grade of this material on the overall gold mill head grade during Years 3 through 7.

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Table 16-5: Proposed Mine Production Schedule

Material	Pre-Prod.	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9	Year 10	Year 11	Year 12	Year 13	Year 14	Year 15	Year 16	TOTAL
Cutoff nsr $/t	15.00	17.50	15.00	11.00	12.00	12.00	12.00	12.00	9.20	9.20	9.20	9.20	9.20	9.20	9.20	9.20
DIRECT MILL FEED: Mine To Mill: Indicated
ktonnes	4,234	10,639	18,007	17,034	17,820	18,463	18,855	18,152	16,023	14,583	11,883	12,788	13,520	14,079	14,884	13,822	2,580	237,366
NSR, $/t	36.70	36.96	36.87	38.77	29.28	31.21	28.56	24.10	26.49	27.79	35.67	40.63	33.75	28.15	24.75	27.98	48.91	31.40
Au, g/t	1.034	1.048	1.063	1.134	0.847	0.907	0.833	0.700	0.778	0.819	1.040	1.191	0.979	0.821	0.715	0.819	1.448	0.913
Ag, g.t	6.59	7.05	6.05	5.06	4.61	4.49	3.77	3.31	2.64	2.45	4.96	5.21	5.49	3.90	3.85	3.11	3.610	4.37
MINERALIZED MATERIAL: Mine To Mill: Inferred
ktonnes	1,549	3,288	3,893	4,866	4,080	3,437	3,045	3,748	4,377	5,817	8,517	7,612	6,880	6,321	5,516	6,186	2,581	81,713
NSR, $/t	32.02	31.36	29.01	23.35	24.50	22.37	21.24	19.66	18.01	24.12	29.02	33.06	23.23	23.04	24.17	26.55	38.16	25.87
Au, g/t	0.896	0.885	0.828	0.671	0.702	0.642	0.612	0.568	0.526	0.709	0.841	0.970	0.671	0.665	0.699	0.778	1.133	0.749
Ag, g.t	5.98	5.96	5.22	3.95	4.38	3.89	3.55	3.12	2.20	2.47	4.63	4.02	3.96	3.87	3.77	2.82	2.550	3.83
Low Grade to Stockpile: Indicated
ktonnes	701	5,606	4,195	584	1,408	1,850	2,100	2,125										18,569
NSR, $/t	12.59	13.90	12.78	10.50	11.07	10.96	11.05	11.05										12.33
Au, g/t	0.344	0.377	0.349	0.285	0.309	0.310	0.316	0.315										0.341
Ag, g.t	3.24	4.23	3.86	3.15	2.56	2.34	2.05	2.05										3.26
Low Grade to Stockpile: Inferred
ktonnes	501	2,989	1,605	364	560	699	844	742										8,304
NSR, $/t	12.85	13.67	12.73	10.51	11.07	11.09	11.01	11.08										12.41
Au, g/t	0.348	0.373	0.350	0.290	0.303	0.313	0.314	0.319										0.343
Ag, g.t	3.53	3.88	3.59	2.78	3.03	2.45	2.20	1.83										3.22
Hi Silver to Stockpile: Indicated
ktonnes				0	0	0	0	0										0
NSR, $/t
Au, g/t
Ag, g.t
Hi Silver to Stockpile: Inferred
ktonnes				80	2,107	5,241	2,383	164										9,975
NSR, $/t				15.17	16.31	17.48	15.45	11.18										16.63
Au, g/t				0.075	0.078	0.070	0.062	0.027										0.069
Ag, g.t				35.82	41.41	48.47	43.72	34.01										45.50

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

16.3 Waste Management

Five types of waste rock are currently classified at Blackwater:  overburden, two types of non-acid generating rock (NAG4 and NAG3), and two types of potentially acid generating rock (PAG2 and PAG1).  The production of waste rock over the LOM is indicated in Table 16-6.  Some of the material types will be used for the construction of the tailings storage facility (TSF) embankments during the mine life, and the rest will be stored in various locations near the planned pit, as shown in Table 16-7 and Figure 16-3.

The waste dumps and stockpiles have been designed to hold all waste materials and low-grade mill feed tonnages not delivered to the primary crusher or the TSF facility.

The overburden dump is east of the pit, and NAG3 material is placed on top of the overburden.  Each dump will progress from south to north through the mine schedule.  The overburden provides a base for the NAG3 material, as discussed in Section 15.  The top of the overburden dump will have a slight pitch toward the open pit to direct any runoff to the pit.  The NAG4 dump is west of the open pit.

A low-grade stockpile will be north of the pit and will be segregated between low-grade mill feed and high-silver mill feed.  The stockpile will be created during the preproduction year through Year 7.  The high-silver mill feed is removed from the stockpile during Years 8 through 14 and the low-grade mill feed is removed during Years 14 through 17.

The design of the waste dumps and stockpiles used dry bulk densities of 1.48 tonnes per cubic metre (t/m3) for overburden and 2.03 t/m3 for all rock materials.  This assumes a 30% swell from pit in-place volume to volume in the dump or stockpile.

The maximum tonnage and volumes of the dumps and stockpiles are shown in Table 16-8.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Table 16-6: Waste Tonnage

Total Waste Tonnage
ktonnes	Pre-Prod.	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9	Year 10	Year 11	Year 12	Year 13	Year 14	Year 15	Year 16	Year 17	TOTAL
overburden	12,263	9,881	5,857	17,314	9,038	768	6,698	9,111	13,532	11,681	2,764	0	0	0	0	0	0	-	98,907
NAG4	1,971	4,099	3,772	9,967	7,512	2,739	8,785	13,210	17,823	23,577	21,137	8,293	5,007	1,994	1,953	838	0	-	132,677
NAG3	2,025	3,107	3,107	4,185	10,001	9,135	6,531	7,345	5,077	5,443	10,459	7,625	7,744	3,084	2,025	2,236	189	-	89,318
PAG2	1,906	4,020	3,217	7,418	6,281	7,018	9,611	10,052	9,064	8,459	10,471	8,340	7,117	3,104	1,821	1,528	202	-	99,629
PAG1	4,850	16,371	16,347	35,188	38,193	47,650	38,148	32,351	31,104	27,440	29,351	17,393	26,631	20,212	15,765	18,561	3,016	-	418,571
Total	23,015	37,478	32,300	74,072	71,025	67,310	69,773	72,069	76,600	76,600	74,182	41,651	46,499	28,394	21,564	23,163	3,407	0	839,102
Overburden, Total	12,263	9,881	5,857	17,314	9,038	768	6,698	9,111	13,532	11,681	2,764	0	0	0	0	0	0	-	98,907
Pit to Embankments	860	3,122	990	7,917	9,038	768	874	794	582	614	643	-	-	-	-	-	-	-	26,202
Pit to Dumps	11,403	6,759	4,867	9,397	0	0	5,824	8,317	12,950	11,067	2,121	-	-	-	-	-	-	-	72,705
Dump to Embankment	-	-	-	-	-	-	-	-	-	-	-	465	585	370	378	390	549	-	2,737
NAG4, Total	1,971	4,099	3,772	9,967	7,512	2,739	8,785	13,210	17,823	23,577	21,137	8,293	5,007	1,994	1,953	838	0	-	132,677
to Embankments	1,363	1,923	2,129	9,967	7,512	2,739	8,785	10,989	10,085	13,283	2,622	1,787	1,448	858	648	434	-	-	76,572
to Dumps	608	2,176	1,643	0	0	0	0	2,221	7,738	10,294	18,515	6,506	3,559	1,136	1,305	404	-	-	56,105
Dump to Embankment	-	-	-	-	-	-	-	-	-	-	-	-	-	-	-	-	219	-	219
NAG3, Total	2,025	3,107	3,107	4,185	10,001	9,135	6,531	7,345	5,077	5,443	10,459	7,625	7,744	3,084	2,025	2,236	189	-	89,318
to Embankments	2,025	393	3,107	4,185	4,252	4,609	4,012	3,208	2,857	3,090	2,727	2,271	2,328	1,968	1,873	1,968	189	-	45,062
to Dumps	0	2,714	0	0	5,749	4,526	2,519	4,137	2,220	2,353	7,732	5,354	5,416	1,116	152	268	0	-	44,256
Dump to Embankment	-	-	-	-	-	-	-	-	-	-	-	-	-	-	-	-	2511	-	2,511
PAG2, Total	1,906	4,020	3,217	7,418	6,281	7,018	9,611	10,052	9,064	8,459	10,471	8,340	7,117	3,104	1,821	1,528	202	-	99,629
to Embankments	1,906	4,020	-	-	-	-	-	-	-	-	-	-	-	-	-	-	-	-	5,926
to TSF sub-aqueous	0	0	3,217	7,418	6,281	7,018	9,611	10,052	9,064	8,459	10,471	8,340	7,117	3,104	1,821	1,528	202	-	93,703
PAG1, Total	4,850	16,371	16,347	35,188	38,193	47,650	38,148	32,351	31,104	27,440	29,351	17,393	26,631	20,212	15,765	18,561	3,016	-	418,571
to Embankments	4,850	1,897	-	-	-	-	-	-	-	-	-	-	-	-	-	-	-	-	6,747
to TSF sub-aqueous	0	14,474	16,347	35,188	38,193	47,650	38,148	32,351	31,104	27,440	29,351	17,393	26,631	20,212	15,765	18,561	3,016	-	411,824

Table 16-7: Waste Storage Location

Material Type	Use in TSF Construction	Storage of Remaining Tonnage
Overburden	Embankment Core, US & DS mass fills	Stockpile adjacent to Pit
NAG4 rock	US & DS mass fills	Stockpile adjacent to Pit
NAG3 rock	US only mass fills	Stockpile on top of overburden stockpile with runoff into pit
PAG2 rock	Dam C only, US & DS fills	Sub-aqueous storage inside TSF pond
PAG1 rock	Dam C only, US & DS fills	Sub-aqueous storage inside TSF pond

Note:  US = upstream side of the embankment core.  DS = downstream side of the embankment core

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Table 16-8:  Maximum Waste Dump and Stockpile Volumes and Tonnages

Dump	Volume (m3 x 1000)	ktonnes
Overburden	50,974	75,442
NAG3	21,801	44,256
NAG4	27,746	56,324
Stockpiles:
Low Grade Mill Feed:	13,238	26,873
High Silver Mill Feed:	4,914	9,975

16.4 Geotechnical Parameters

The Project site is characterized by the following sequence of overburden and bedrock:

· Post-glacial Fluvial and Organic Materials – The valley bottoms in the Project area contain mainly fluvial deposits.  Most of the modern streams are meandering, with channels consisting primarily of fluvial sands and gravels.  The floodplains of these streams are deposits of fine sands, silts, and organic material.  In the flat, open areas on the valley floors, the landscape is dominated by marshes and shallow lakes filled with organic sediments formed from decaying marsh vegetation.  Large accumulations of peat can be found in areas where drainage was restricted during the post-glacial period.

· Outwash Fluvial Deposits – These deposits were formed at the head of the advancing and retreating ice sheet.  They are typically found as dry, compact, sand, gravel and cobble materials forming a terrace on either side of Davidson Creek between 10 and 18 m thick overlying glacial till deposits.

· Esker Deposits – A series of meltwater channels extend north as eskers in the Chedakuz Valley.  Another large esker complex is found on the western margin of Top Lake Valley as it cuts through the Fawnie Range.  These eskers fan out into the Chedakuz Valley, which suggests that they were formed sub-glacially as meltwater flowed out of the Top Lake Valley.

· Glacial Till Deposits – Glacial till deposits are the most dominant Quaternary deposits in the region.  Deposit thickness can be quite variable, ranging from a few to several metres.  The deposits are generally very dense with low permeability, and consist of well-graded, silty sandy gravel.  The contact between the basal till and the underlying bedrock tends to be quite clear and well defined in most areas.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

· Glaciofluvial Deposits – Typically these water-bearing sand and gravel layers are interbedded within sequences of glacial till and were deposited during inter-glacial ice melt periods.  These layers are typically deposited as non-continuous lenses 3 m to 5 m thick.  The distribution and prevalence of these layers will be investigated further throughout the 2012 site investigation programs.

· Glaciolacustrine Deposits – These deposits typically consist of horizontal laminated fine sands, silts, and clays that were deposited in glacial lakes during inter-glacial ice melt periods.  Glaciolacustrine deposits interbedded in sequences of glacial till are thought to be uncommon and will be investigated further during the 2012 site investigation programs.

· Ootsa Andesite Bedrock – Andesite bedrock is typically found within 10 m depth on topographic highs and up to 85 m depth in valley bottoms.  Bedrock was typically found to have the following geotechnical properties:

- Highly weathered for up to the first 20 m, then becoming more competent and moderately to slightly weathered.

- Rock Mass Rating classification (RMR 89) of “fair” to “good,” with values ranging from 40 to 80 with a mean of 65.

- Rock Quality Designation (RQD) values ranging from 30% to 90%, with a mean of 70%.

- Low permeability with hydraulic conductivity ranging from 10-6 to 10-4 cm/s.

- Rock strength of “medium strong” to “strong,” with unconfined compressive strength (UCS) laboratory test results ranging from 37 to 193 MPa with a mean of 97 MPa, Young modulus 23 to 68 GPa, and Poisson ratio of 0.20 to 0.25.

16.4.1 Proposed Open Pit Area

The proposed open pit will reach a maximum depth of 400 m.  Knight Piésold Ltd. completed a desktop-based preliminary pit slope geotechnical assessment for the Project in 2011 and a geomechanical drilling program in early 2012.  Preliminary pit slope designs were completed to establish allowable overall slope angles.

The bedrock at the Blackwater deposit consists of intercalated volcanic and volcaniclastic felsic to intermediate lapilli and ash tuff, volcanic breccia, and andesitic flows.  The strata are laterally discontinuous over short distances and generally dip to the northwest.  The bedrock is pervasively hydro-fractured and silicified with extensive broken zones within the deposit area.  The deposit is bounded by sub-vertical faults to the west and north.  The north and west extent of gold mineralization is yet to be delineated.  There is evidence of large-scale faults, shear zones, and predominant structures in the rock mass.

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Four major geotechnical domains were defined for a preliminary pit slope geotechnical assessment:

· Overburden (OB)

· Felsic Volcanics (FV)

· Andesite (AND)

· Broken Zone (BZ).

The overburden thickness typically ranges from 5 m to 10 m, but increases to 50 m to 80 m along the east side of the deposit.  The intact strength of all rock types was found to be “medium strong” to “strong,” with average UCS values between 50 and 180 MPa.  The rock mass quality appears to be “very poor” for the broken zone unit, with an average RQD value of 15.  The rock mass quality for the rest of geological units was characterized as “fair” to “good.” with an average RQD value of over 50.

Groundwater levels were near surface in the lower elevation areas of the deposit.  Artesian flows were encountered in some drill holes on lower slopes.  The groundwater table was found at approximately 35 m depth in drill holes on the higher-elevation south side of the deposit.  Given the presence of broken rock in the deposit area, the rock mass permeability is expected to be high.

Preliminary pit slope recommendations are based on a review of the currently available geomechanical information, stability analyses results, assumptions on mining equipment, and experience at similar open pit operations in the Project region.  The open pit area has been divided into design sectors with preliminary overall slope angles as shown in Figure 16-4.

Drilling along the proposed south and east pit walls to support more detailed studies was completed in the spring of 2012.  The second phase of the program, which targets the west and north pit walls, is planned to be completed in late 2012.

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October 2012

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

16.4.2 Proposed Tailings Storage Facility Areas

Two sites, designated Sites C and D, are proposed to be utilized for tailings storage.

The Site D TSF embankment is located to optimize the natural topography of the Davidson Creek Valley to allow for efficient and long-term storage of tailings and waste rock materials that will arise from mining activities in the proposed open pit.  The area is characterized by thick glacial till deposits 15 m to 70 m thick overlying bedrock.  Initial investigations found inter-bedded glaciofluvial sand and gravel layers within the glacial till.  These are potential water seepage pathways beneath the TSF.

Sand and gravel glacial outwash deposits overlie the glacial till to form a terrace between 10 m and 18 m thick on either side of Davidson Creek.  These deposits are typically very permeable.

A large esker deposit, which flowed within and under an ice sheet, overlies the glacial till deposits to the north of the proposed TSF site.  The bedrock geology consists of andesite volcanic rocks from the Cenozic aged Oosta Lake Formation at depths ranging from 15 m to 70 m.  Bedrock was found to be highly weathered for the first 20 m, followed by moderately to slightly weathered rock.  In-situ hydraulic conductivity testing has shown the rock mass is generally of low permeability.

The Site C TSF embankment foundation conditions are characterized by glacial till sequences with inter-bedded discontinuous sand and gravel deposits.  The bedrock geology consists of andesite volcanic rocks from the Cenozic aged Oosta Lake Formation at depths ranging from 40 m to 85 m.  Bedrock was found to be highly weathered for the first 25 m, followed by moderately to slightly weathered rock.  In-situ hydraulic conductivity testing has shown the rock mass is generally of low permeability.

Organic peat deposits found in valley bottoms will need to be stripped before embankment construction for the TSF.  These deposits and other unsuitable overburden material will be stockpiled in stable configurations and retained for use in reclamation and closure activities.

16.4.3 Proposed Plant Site Area

The proposed plant site is located on a topographic high between the open pit and the TSF site at an approximate elevation of 1,425 masl.  The surface consists of approximately 6 m of compact to dense sand and gravel deposits overlying competent andesite volcanic rocks of the Oosta Lake Formation.  The water table was encountered at 5 m depth in the sand and gravel unit.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Future investigation programs will investigate the extent, depth, and occurrence of the sand and gravel layers.  The excavated sand and gravel may be suitable for use in concrete production or as backfill material.

16.4.4 Aggregate Potential

Potential aggregate sources for construction of the Project facilities at the site consist of the esker complex north of the TSF, sand and gravel from the cut-off trench excavation at the TSF Site D embankment, and/or surplus rock excavation from plant site grading.  Laboratory testing will be conducted to confirm the suitability of these materials for concrete aggregate.

16.5 Hydrogeology

Open pit dewatering will continue throughout the mine life, with dewatering flows either directed to the TSF or discharged directly to the process water tank at the mill site.  Once active mining from the pit has ceased, pit dewatering will cease and some surface water will be diverted to the open pit to contribute to pit filling.  TSF supernatant will also be pumped to the open pit, which will allow for natural degradation of contaminants in the TSF water.

16.6 Mining Equipment

Mine equipment requirements were calculated based on the annual mine production schedule, the mine work schedule, and equipment shift production estimates.  The size and type of mining equipment is consistent with the size of the Project: peak run-of-mine material movements of 97 million tonnes per year.  IMC determined the optimum production schedule, and then calculated the mining fleet requirements.

There is sufficient equipment to perform the following duties:

· Construct additional roads, after preproduction, as needed to support mining activity, including pioneering work necessary for mine and dump expansion

· Strip topsoil in advance of mining and dumping

· Mine and transport the mill feed to the crusher (or crusher stockpile); mine and transport the waste material from the pit areas to the waste storage areas and the TSF construction areas

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

· Doze and compact with haul trucks the upstream and downstream portion of the TSF embankments as waste tonnages are delivered by the mine truck fleet to the TSF

· Deliver overburden by the mine fleet to the TSF for construction of the compacted till core by a contractor

· Maintain all the mine work areas, in-pit haul roads, waste storage areas, crusher stockpiles, and external haul roads

· Build and maintain in-pit and on-dump drainage structures as required.

A summary of the total fleet by year for the mine major equipment is shown in Table 16-9.  Table 16-10 illustrates the number of support equipment units that will be required.

The mine work schedule, the scheduled shifts per year, and the expected shifts available after accounting for holidays are summarized below:

· The mine is scheduled to work two shifts per day, 365 days per year.

· The mine will operate seven days per week, two 12 hour shifts per day.

· There are 20 planned shifts lost to weather leaving a total of 710 available shifts each year.

· Four mining crews are necessary to cover the required shifts throughout the mine life.

16.6.1 Operating Time per Shift

Operating time per shift represents the actual time during the shift that the equipment is productive.  This is equal to the total scheduled time minus all scheduled and unscheduled delays.  For a 12-hour shift, the equipment will accumulate 575 minutes (9.6 hours) of metered time.  Delays are caused by shift changes, lunch and breaks, blasting, equipment inspections, and fuelling equipment.  About 479 minutes (8 hours) will be the net productive operating time, based on a job efficiency of 83.3% (50 minutes per hour).  The job efficiency factor accounts for unscheduled delays in the operation due to interruptions caused by other mine activities and non-productive use of time by operators and supervisors.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

16.6.2 Material Characteristics

The material characteristics used for the equipment productivity calculations have been divided into three basic rock type categories for this study: mill feed, waste rock, and overburden.  The materials were divided to account for the different densities for the equipment fleet calculations.  Dry bank density is 2.74 t/m3 for mill feed and waste rock and 2.00 t/m3 for overburden.

Moisture content of 3.0% was used for all material types.  The material-handling swell for equipment was estimated at 40% for run-of-mine material.  This is greater than the swell used for waste storage area volumes, which is 30%.  Some compaction is expected in the waste storage areas.  Moisture content represents weight percent of the dry weight of the material.  The density of wet loose material (2.02 for mill feed and waste rock, 1.47 for overburden) was used to calculate haul truck speeds and allowable payload limits.

All equipment production is reported in dry metric tons.

16.7 Mine Personnel

Mine personnel include all salaried supervisory and staff working in the mine operations, maintenance, and engineering/geology departments, and the hourly personnel required to operate and maintain the drilling, blasting, loading, hauling, and mine support activities.  In general mining activities end once the mill feed is delivered to the crusher.

16.7.1 Salaried Staff

Mine salaried staff requirements over the Project life are shown in Table 16-11.  The staff builds to a total of 40 by Year 1 and reduces to 38 in Year 3, when one operations trainer and one maintenance trainer are retired.

16.7.2 Hourly Personnel

Mine total hourly personnel requirements are shown in Table 16-12.  The required number of personnel is 163 by the end of preproduction.  The maximum number during commercial production is 421 in Year 9 of production.

Most personnel in mine operations are equipment operators.  The number of operators was based on the major equipment fleet requirements.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

The fleet size, the number of crews, and the fleet utilization are used to calculate the number of operators as follows:

Calculated number of Operators = (Total Fleet) x (Crews) x (Utilization) x (Cross Training Factor)

The operator cross-training factor (CTF) was assumed at 1.31 for this study.  This factor accounts for shifts when more than the average number of units are available and need to be scheduled and for inefficiencies in cross-training operators.  The limits on this factor are 1 (100% cross-training of

operators) to 1 divided by the utilization, which implies no cross-training.  The result of no cross-training in the above calculation of operators is that every piece of equipment is manned on every shift (total fleet x four crews).  The factor of 1.31 is based on minimal cross-training of operators.  Where the calculated number of operators had a fraction of 0.2 or more, the assigned number of operators was rounded up.

16.8 Comments on Section 16

In the opinion of the QPs, the following conclusions and comments are appropriate to the proposed mine plan:

· The open pit mining method will use conventional mining techniques and be based on a truck-and-shovel operation.

· The mine plan includes both Indicated and Inferred Resources to make up the mill feed tonnage.  Material not processed is waste.

· The LOM production will be 255.9 million tonnes (Mt) from the Indicated category and 100.0 Mt from the Inferred category.   Mill feed will be delivered to a primary crusher that feeds a whole ore leach (WOL) gold–silver recovery plant.  The milling rate is 60,000 metric tonnes (tonnes) per day (t/d), or 21.9 Mt per year (Mt/a).  The mine life is 16.4 years after commercial production begins.  The total tonnage mined from the open pit is 1,195 Mt at an initial daily mining rate of 169,000 t/d and increasing to 273,000 t/d during Years 3 through 10 before declining after the peak waste mining period.  A total of 7.2 million ounces of gold and 35.3 million ounces of silver are estimated to be contained in the Indicated category of mill feed tonnage during the life of the mine.  An additional 2.1 million ounces of gold and 25.5 million ounces of silver are estimated to be contained in the Inferred category of mill feed.

· Some of the mill feed, referred to a low-grade mineralized material, will be stockpiled in the early years for re-handle into the mill at the end of the mine life.  Another sub-set of the mill feed is within a high silver–low gold area in the northern part of the deposit, referred to as high silver mill feed.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

· The proposed open pit will be mined in four phases.  Pit design has appropriately considered available geotechnical information.

· Five types of waste rock have been identified and the mine design appropriately incorporates consideration of these materials.  The waste storage facilities are designed to accommodate the waste rock production over the LOM.

· Equipment and personnel requirements have been considered in relation to the mine plan and are appropriate for the mine production rate, sequencing, and mine life.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

17.0 RECOVERY METHODS

17.1 Proposed Process Flow Sheet

The whole ore leach or WOL plant is based on conventional technology and proven equipment.

Run-of-mine mineralized material from the open pit will be crushed and conveyed to the concentrator where the mineralization will be ground to allow gold extraction and recovery.  Some gravity-recoverable gold will be recovered from processing a bleed stream within the grinding circuit.

During initial operations, the plant will be processing primarily oxide and transition materials, which are softer, and therefore throughputs as high as 70,000 t/d of transition material and 85,000 t/d of oxide material could potentially be achieved.

Mineralization ground to a P80 of 150 µm will be thickened ahead of leaching to optimize slurry density and leached in a conventional cyanide leach circuit.  Gold will be recovered from loaded carbon, electrowon, and smelted into doré bullion.  CIP residue will be thickened and treated in the cyanide destruction circuit prior to disposal in the TSF.

The proposed process flow diagram for WOL is shown in Figure 17-1.

17.2 Plant Design

Selected process design criteria for the WOL flowsheet are summarized in Table 17-1.

The process is based on conventional technology, with equipment well-suited for the application and material processed.  The selected process option accommodates the sequential processing of ores with substantially differing physical characteristics while keeping the plant as simple as possible.

The mine schedule generates predominately oxide and transition mineralized material feed in the first two years of operation, followed by almost entirely sulphide material for the remaining mine life.  There is an opportunity to increase mill throughput during the first two years because of the softer nature of oxide and transition materials; this will be investigated during more detailed studies.

Grinding, flotation, regrind, gold desorption, and the gold room will be enclosed in the process building.  Other unit operations—cyanide leaching, cyanide destruction, and thickening—will be located outside the building in bunded, contained areas.

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Table 17-1:  Process Design Criteria - WOL Flowsheet

17.2.1 Crushing and Coarse Ore Stockpile

The primary crushing and conveying area includes a crusher station near the pit rim, the overland conveyor, and the coarse ore stockpile (COS).  Major equipment will include a primary crushing station complete with a truck dump hopper, gyratory crusher, support structure, rock breaker, maintenance crane, apron feeder, and overland conveyor to transport mined material to the coarse ore stockpile.

A 1.52 m x 2.26 m (60" x 89") primary crusher will operate at 3,330 t/h and crush mineralized material to 80% passing 125 mm.  The crushed product will flow by gravity to a discharge pocket from where it will be fed by an apron feeder at a controlled rate to the coarse ore stockpile feed conveyor.

The coarse ore stockpile will have a live load of 60,000 t, equivalent to 24 hours’ supply of mineralized material to the mill; the total live plus dead capacity will be approximately 350,000 t.  The stockpile will not be covered.  Coarse mineralized material will be reclaimed at rates determined by the SAG mill demand.  A track dozer will be used to push mineralized material into the feeders when upstream maintenance activities require mineralized material to be reclaimed from the dead section of the stockpile.  The reclaimed mineralized material will be transferred onto the SAG mill feed conveyors.

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Each SAG mill feed conveyor will run in a separate reclaim tunnel beneath the coarse ore stockpile.  The three apron feeders for each conveyor will be supported within the concrete tunnel.  The SAG mill feed conveyors will extend from below the coarse ore stockpile to the SAG mills, and will transition from a tunnel to a uncovered section with a walkway and handrails on both side until they enter the concentrator.

17.2.2 Grinding

The grinding circuit will consist of a two 11.0 m x 6.7 m effective grinding length (EGL) SAG mills in open circuit followed by two 12.6 m x 19.3 m EGL ball mills in closed circuit with cyclones.  Pebbles – material greater than 20 mm leaving the SAG mills – will be screened at the SAG mills discharge and conveyed to one of two large pebble crushers, which will discharge the product onto the SAG mill feed conveyor for recycling into the SAG mill.  The average product size from the grinding circuit will be 80% passing 150 μm. The dual- pinion SAG mills will be powered by two 8 MW motors.  The ball mills will each be powered by dual 8 MW drives.

For each SAG mill, discharge with a top size of 90 mm will pass directly onto a vibrating screen equipped with a 20 mm screen deck.  Undersize from the vibrating screen will collect in the SAG and ball mill discharge pumpbox.

The washed pebbles in the screen oversize from each circuit will drop onto the pebble transfer conveyor, which discharges onto the pebble feed conveyor for delivery from the grinding bay to the pebble crushers.  The pebble crushers will be in a separate building between the mill building and the COS.  In addition to the crushers, the assembly will consist of a surge bin, bypass chute, tramp metal detection removal magnets, and pebble crusher feeders, complete with access stairways and platforms.  The crushers will discharge onto the pebble discharge conveyor, which in turn will return the material to the SAG mill feed conveyor.

Pebbles from the pebble crusher feed conveyor will enter the crusher feed surge bin and be metered into one of two pebble crushers by a belt feeder.  When the crushers are not available, or if metal is detected on the feed conveyor to the surge bin, or when pebble production is excessive, pebbles will be diverted through a bypass chute to the SAG mill feed conveyor.  The pebble crushers will be 1,000 kW units and will reduce the material to a P80 of 13 mm.

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17.2.3 Gravity Separation

Approximately 15% of the cyclone underflow mass will be bled off to the gravity scalping screen to remove +2 mm material, which will report to the ball mill feed hopper, while the screen undersize will be processed through a centrifugal gravity concentrator to recover free liberated gold.  Gravity tailings will be directed to the mill feed hopper, and the free gold concentrate will be discharged into a storage hopper from where it will be treated in an intensive cyanidation unit (ICU).  The small volume of pregnant solution from the ICU will be blended in with the main plant pregnant eluate.  ICU residue will be pumped back to the grinding circuit.

This circuit is expected to recover approximately 12% of the total gold.  This estimate, and optimal sizing of the gravity circuit, will be verified in future testwork.

17.2.4 Leaching and CIP

Primary cyclone overflow will be screened to remove trash and directed to a leach conditioning tank, where lime will be added to increase the pH to a range of 10.5 to 11 in preparation for cyanide leaching.  The cyclone overflow will report to a feed leach thickener, and thickener underflow at 50% solids will report to the leaching circuit.

The leach circuit will consist of three trains of eight 18 m diameter leach tanks.  The slurry will be split between the trains to achieve a leach residence time of 30 hours.  Cyanide will be added in stages to the first leach tank to optimize gold extraction and cyanide consumption.

The discharge from the leach trains will be recombined and directed to six 18 m diameter CIP tanks for gold and silver adsorption.

17.2.5 Carbon Desorption and Regeneration and Gold Recovery

Loaded carbon from both the WOL and concentrate leach CIP circuits will be pumped into a 20 tonne capacity stripping vessel.

Pressure AARL (Anglo American Research Laboratory) has been selected for the carbon stripping process, where the carbon is acid-washed in dilute hydrochloric acid, water-washed, then soaked in 3% sodium cyanide and a 1% to 2% sodium hydroxide solution for approximately 30 minutes.  The carbon is eluted with a 6- to 10-bed volume of deionized water at a temperature of 110° to 120°C and pressure of 70 to 100 kPa.  Elution is estimated to be complete in 12 hours.

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After elution, carbon will be regenerated in an electric regeneration kiln.  Once quenched and screened, the reactivated carbon will be fed back into the last leach tank.

Eluated solution will be pumped to electrowinning cells to produce a gold-silver precipitate sludge.  Fully loaded cathodes will be periodically removed from the cell, and gold/silver sludge will be washed off with high-pressure water.  The sludge will be calcined and then mixed with fluxes and smelted on site to produce gold-silver doré.

17.2.6 Tailings Thickening

Slurry discharging from the CIP tanks will report to an 80 m diameter CIP residue thickener prior to cyanide destruction.  This thickener will recover some of the cyanide-bearing solution for recycle to the leaching process, which will aid in reducing both cyanide consumption and cyanide destruction costs.  Slurry will be thickened to 55% solids before cyanide destruction.  The thickener overflow will report to a leach process water tank for recycling.  Solution balance in the plant is maintained by maximizing the residue thickener density relative to feed leach thickener underflow.

17.2.7 Cyanide Destruction

The cyanide destruction circuit will use a typical one-stage SO2-Air process.  The process removes cyanide (including precipitating ferricyanide) contained in the tailings slurry and is considered to be effective in achieving residual concentrations of less than 1.0 mg/L Weak Acid Dissociable Cyanide (WAD) and 2 mg/L total CN.

The system will consist of a one-stage reactor, SO2 storage and feed system, lime or caustic chemical feed system, copper sulphate feed system, and aeration system.

17.3 Energy, Water, and Process Materials Requirements

17.3.1 Reagents

Table 17-2 lists the process reagents, their purpose, and consumption levels.

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Table 17-2: Reagent List

Reagent	Circuit	Main Purpose	Consumption kt/a
Cyanide	Leaching	Lixiviant for gold and silver	22.0
Lime	Leaching	pH modifier to minimize cyanide loss	30.6
Flocculants	Thickening	-	0.4
Carbon	CIP	Gold and silver adsorption	1.7
Copper Sulphate	Cyanide destruction	-	1.4
SO2	Cyanide destruction	Cyanide destruction	9.0

17.3.2 Energy and Water

These requirements are discussed in Section 18 (water) and Section 21 (power).

17.3.3 Cyanide

The Project is expected to consume approximately 22,000 t/a, or 1,825 t/month, of sodium cyanide.  All transport, unloading, and storage will be accordance with the International Cyanide Management Code (the Code, or ICC) of the International Cyanide Management Institute (ICMI).  The ICMI is a voluntary program for the gold mining industry that promotes responsible management of cyanide to enhance the protection of human health and reduce the potential for environmental impacts.

New Gold will select an ICMI-certified supplier, who will ensure that all personnel handling and managing cyanide transportation are trained and experienced in ICC procedures.

For the purposes of the PEA, it is assumed that cyanide will be supplied in bulk, preferably in solid briquette form, from a manufacturing facility in the USA.  Delivery will likely be by railcars of 80 tonne capacity to a secure B.C. regional terminal facility near the mine access road, where the cyanide will be transloaded from the railcars into Isocontainers of approximately 20 tonne capacity and trucked to the mine.  The terminal will be operated under the supervision of the cyanide supplier and will be designed and operated according to the ICC.

Isocontainers are double-walled tanks that provide a greater degree of environmental security.  At the plant site, the full Isocontainers will be off-loaded into secure storage for use in the cyanide make-up facilities.  Trained operators will mix the solid cyanide with water for dissolution to a predetermined concentration that will be metered directly into the process use areas.  Empty Isocontainers will be backhauled to the terminal facility for reuse.

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At the anticipated consumption rates, about 20 railcar deliveries will be made to the terminal facility and 75 Isocontainers hauled to the mine site per month.  Some inventory storage is likely to be maintained both on railcars at the terminal ‒ enough for about two months’ use ‒ and in Isocontainers at the mine site ‒ enough for two to four days’ use.

Alternatively cyanide may be supplied from a manufacturing facility outside of North America.  Delivery would likely be in 1 tonne bag/boxes packed in 20 tonne Isocontainers.  The Isocontainers would be railed to a transloading facility and trucked from there to the mine.  The full containers would be off-loaded and de-stuffed at the mine, and the bags will be manually unloaded into the cyanide solution make-up system.  The used box and washed bag packaging will be incinerated at site under strictly controlled conditions and the containers returned to the terminal.

17.4 Comments on Section 17

In the opinion of the QPs, the following conclusions and comments are appropriate:

· The Project will use conventional mineral processing equipment to produce doré.  The plant design is based on conventional technology and proven equipment.

· The process design is based on the metallurgical testwork and is appropriate to the grind, flotation and recovery characteristics defined for the different mineralization types.

· Run-of-mine mineralized material from the open pit will be crushed and conveyed to the concentrator where the material will be ground to allow gold extraction and recovery.  Some gravity-recoverable gold will be recovered from processing a bleed stream within the grinding circuit.

· During initial operations, the plant will be processing primarily oxide and transition materials, which are softer, and therefore throughputs as high as 70,000 t/d of transition mineralization and 85,000 t/d of oxide material could potentially be achieved.

· Mineralized material ground to a P80 of 150 µm will be thickened ahead of leaching to optimize slurry density and leached in a conventional cyanide leach circuit.  Gold will be recovered from loaded carbon, electrowon, and smelted into doré bullion.  CIP residue will be thickened and treated in the cyanide destruction circuit prior to disposal in the TSF.

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The WOL process configuration was selected for the current PEA based on the available metallurgical data as of May 2012.  The WOL, FCL, and Hybrid process options need further investigation to confirm which is optimal from an economic, operational, and environmental perspective.  A program of additional metallurgical testwork, which AMEC has helped design, is underway to test the variability and overall metallurgical performance of these alternative flowsheets.  In particular, the following aspects should be studied:

· WOL testwork at lower cyanide concentrations to determine if a more economically favourable combination of leach time and concentration results in lower cyanide consumption.

· Flotation testwork on additional oxide and transition samples, including blends with sulphide material, to confirm if acceptable gold recoveries can be achieved while reducing the capital currently allocated to leach tanks for the Hybrid circuit WOL mode.

· Concentrate leach testwork to economically optimize the combination of leach time, regrind size, and cyanide concentration.

· Gravity-recoverable gold (GRG) characterization to predict gravity gold recoveries at a plant scale and to infer the magnitude of the overall gold recovery benefit in all milling circuit configurations.

· Mineralogy and gold deportment.

· Metallurgical response of low-grade gold and high-grade silver stockpile materials.

· Continue to monitor and review the quality of laboratory data and enhance as required the quality control program, inter-laboratory check assays, and supervision of the metallurgical program.

Trade-off studies within the PEA WOL flowsheet as described in this Report should be investigated to minimize capital and operating costs:

· Evaluate the potential advantages of a carousel-type CIP system versus the conventional CIP.

Trade-off studies should be investigated to minimize capital cost while addressing the potentially challenging flotation performance on oxide and transition mineralized material:

· Evaluate the implications and viability of stockpiling oxide and transition mineralized materials for processing by heap leach, including environmental and economic considerations.

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· Evaluate flotation response of transition and oxide mineralized materials and for stockpiled material for a flotation-only project.

· Optimize the configuration and size of the WOL.

· Re-evaluate throughput selection in light of potential increases in the resource tonnage

· Investigate the possibility of higher throughput for softer oxide and transition materials.

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18.0 PROJECT INFRASTRUCTURE

18.1 Road and Logistics

A study was conducted to identify potential routes to the Blackwater site as well as the current road permit holders, current and future road use, known social, economic, and safety constraints to heavy industrial usage, and potential upgrades required to ensure year-round industrial traffic use.

The primary focus of the report was road transportation options using existing infrastructure where feasible (Allnorth, 2011).  A Design Basis Memo was produced detailing criteria to be used for the review, analysis, and inspection of various ground access routes connecting the Highway 16 corridor south to the Project.

Five potential routes were evaluated.  The selected option for the PEA follows the existing Kluskus Forest Service Road (FSR) that branches south from Highway 16 at Engen.  Figure 18-1 shows the selected PEA road option.

Local forest licensees are and will continue to be the principal users of the roads accessing the Project.

18.2 On-Site Project Infrastructure

The planned on-site infrastructure for the Project includes the following:

· Electrical substation

· Mine haul roads and access roads for plant and site facilities

· Explosives plant/magazine

· Mine truckshop and associated warehouse

· Diesel fuel storage tanks and fuel station

· Site offices

· Warehouse and cold storage

· Medical and emergency response facilities

· Security gate

· Fresh water distribution, potable water treatment plant, and fire water distribution

· Construction laydown area

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· Core logging area

· Microwave communications facility (existing)

· Camp and sewage treatment

· Helipad.

Figure 18-2 shows the preliminary site layout in the context of the open pit, the mine stockpile and dump areas, and the TSF.  Figure 18-3 provides more resolution around the plant, warehouse, administration, and camp facilities.

The primary crusher will be connected by a haul road running directly north of the open pit.  Mine haul roads will continue to the NAG and PAG waste disposal areas to the north and northwest of the pit.  The truckshop, which will be integrated with the warehouse for mine and process spare parts, will be close to the haul road corridor.  The diesel fuel storage tanks will be adjacent to the mine truckshop.

The coarse ore conveyor will continue north from the primary crusher and feed the coarse ore stockpile.  The SAG feed conveyor will connect the coarse ore stockpile to the process plant directly to the east.  The process plant will be located on a hill, which should allow levelling and installation of sound foundations for large equipment such as the SAG and ball mills.  Leach tanks will be positioned along the hill slope to take advantage of gravity flow.

The site electrical substation will be directly beside the plant site.  Administration offices will be south of the process plant and near the substation.

The construction and permanent camps will be constructed northeast of the open pit, and the core storage facility will be west of the camp site.

The site layout allows separation of mine haul traffic from vehicle access to the plant site, administration, and camp facilities.

A construction laydown area will be provided on a ridge between the NAG waste dump and the Site C TSF area.  This area will be used for large equipment, steel, and materials for both major mining equipment and the process plant.

The explosives plant will be sited approximately 2 km west of the pit, away from other site facilities.

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The administration offices will accommodate administration, engineering, geology, mining, and environmental personnel.

The warehouse will be integrated with the truckshop and will stock spare parts for the mine and process plant equipment.

The laboratory will be beside the process plant and will include sample preparation, metallurgical testing, and assay facilities.

18.3 Waste Storage Facilities

18.3.1 Waste Rock Classification

New Gold retained AMEC to conduct a geochemical characterization program of mine materials, including waste rock, mineralized material, overburden, and tailings, for the Project.  From January 2011 to April 2012, a total of 475 samples of waste rock, 41 samples of mineralization, 20 samples of overburden, and 10 samples of tailings were analyzed for acid base accounting (ABA), elemental content, net acid generation testing (NAG test), mineralogical analysis, and leachable metals.  Waste rock and mineralization samples were collected to represent different rock types present at the Project site and also to represent variations in weathering characteristics.  Tailings samples were generated during metallurgical testing of different metallurgical processes that are being considered for the Project (e.g., whole ore leach and/or flotation-leach).

In addition to the samples collected for geochemical characterization, New Gold has compiled a database of over 140,000 metal assays that were derived from analyzing samples over 1 m long intervals of exploration drill core.

Based on the ABA and elemental content data, surrogate parameters for neutralization potential (calcium) and acid potential (sulphur by ICP-MS) have been applied to the extensive exploration database to create a block model for prediction of waste rock and mineralization ABA characteristics.

The following screening criteria were used in the assessment and the block model to identify potentially acid generating (PAG) and non-acid generating (NAG) materials based on the neutralization potential ratio (NPR):

· NPR < 1 Potentially Acid Generating

· 1 < NPR < 2 Uncertain

· NPR > 2 Non-acid Generating

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Results of geochemical block modelling indicate the following:

· The majority of the waste rock is predicted to be PAG or Uncertain (NPR < 2).

· Virtually all of the mineralization is predicted to be PAG or Uncertain.

In addition, the overburden is predicted to be NAG according to ABA.

Most of the waste to be mined from the open pit is expected to be classified as PAG or Uncertain.  However, based on the ABA sample distributions and block modelling results, large blocks of NAG waste are located in the upper east side (overburden) and in the southwest quarter of the open pit.  Additional smaller zones of NAG waste have been identified on the northeastern and northwestern periphery of the pit.  Given their apparent continuity, these blocks may be able to be selectively mined.

In general, metals concentrations in waste rock humidity cell leachate correlate with NPR values in each cell, where the highest leachate metals concentrations were reported by the sample with the lowest NPR value, and the lowest metals concentrations were reported by the sample with the highest NPR value.  Three out of the four waste rock humidity cells with elevated Zn content reported leachate concentrations of Zn that were higher than the Metal Mining Effluent Regulations (MMER) maximum grab sample concentration of 1 mg/L.  Leachate concentrations of Cd were also generally higher than 10 times the BCWQO for all cells.  This suggests that leaching of these metals may be a concern from NAG materials with elevated Cd and Zn contents and from PAG materials before the onset of acidic conditions.  Preliminary evaluation of metal leaching data, including humidity cell and shake flask extraction testing, indicates that waste rock with Zn concentrations below 1,000 ppm may leach lower Zn than samples with higher solid phase Zn concentrations.  Based on this preliminary information, four categories of waste rock were developed for screening waste materials according to their geochemical characteristics for the Project:

· Type 1 (PAG1) PAG rock with NPR <1

· Type 2 (PAG2) Rock with uncertain ARD potential (1<NPR<2)

· Type 3 (NAG3) NAG rock with NPR >2 and solid Zn concentrations >1,000 ppm

· Type 4 (NAG4) NAG rock with NPR >2 and solid Zn concentrations <1,000 ppm.

These criteria were also applied to mineralized material and will be reviewed and refined as additional acid generation potential and metal leaching data become available from samples with lower, more representative Zn content.  A threshold Cd value may also be identified.

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18.3.2 Site Selection

Knight Piésold evaluated locations and methods for mine waste facilities for the Project in April 2012.

18.4 Proposed Tailings Storage Facilities

18.4.1 Tailings Classification

Preliminary ABA testing on tailings samples from the metallurgical testing programs has indicated that the tailings may be PAG.  Additional confirmatory testing is ongoing.

Tailings samples are undergoing testing in humidity cells and report generally low concentrations of dissolved metals, below MMER and BCWQO.

18.4.2 Site Selection

A preliminary assessment of tailings storage facility (TSF) site alternatives was conducted for the Project in January 2011 (Knight Piésold Ltd, January 2011).  Five potential locations for storage of 300 Mm3 of tailings plus potentially acid generating (PAG) waste rock were identified within 10 km of the deposit area.  The alternatives assessment considered economic, operational, and environmental factors and concluded that Sites C and D were the best options due to:

· Relatively high storage efficiency and low embankment volumes

· Close proximity to the mineral deposit area location within the headwaters of Davidson Creek with minimal external upslope catchment area, simplifying operational water management.

The conceptual plans assumed that tailings would be deposited around the perimeter of the TSF from the embankment crests.  The NAG waste was to be disposed of in a waste dump near the pit and/or used to construct the TSF embankments.  The PAG waste material was to be managed using one of two methods:

· Co-disposal within the TSF in such a manner that it becomes progressively saturated by the supernatant pond to prevent oxidation and subsequent acid generation

· Storage in a surface engineered stockpile with a compacted soil liner foundation and cover system.

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The results of this alternatives assessment indicated the preferred mine development concept for future studies should be based on the subaqueous disposal of PAG waste rock in the TSF.

18.4.3 Design Criteria

The design criteria reflect the currently envisaged mine plan and operating strategy.  The conceptual design basis and operating criteria are summarized as follows:

· TSF located on New Gold claims

· IMC production schedule “July 2012 PEA Production Schedule”

· Mill throughput of 60,000 t/d

· Mine life of 17 years

· Whole ore leach tailings generated in the mill process for the first two years to be deposited in TSF Site C, while remaining tailings to be deposited in TSF Site D

· Cyanide leach tailings to be treated with the SO2-Air process prior to deposition in the TSF

· Conventional, thickened tailings disposal

· Overall tailings dry density of 1.3 t/m3

· NAG waste and overburden disposed of in waste dumps near the pit and/or used to construct the TSF embankments

· PAG waste material to be co-disposed of within the TSF in such a manner that it is progressively saturated by the supernatant pond to prevent oxidation and subsequent acid generation

· Waste rock dry density of 2.2 t/m3

· Water for the process sourced from the TSF supernatant ponds

Waste management concepts for the various site materials are as outlined below:

· Topsoil

- temporary stockpiles for use in reclamation

· NAG4/Overburden

- construction material for TSF, where practical and cost-effective

- surplus to NAG4 and overburden dumps

- re-grade and revegetate at closure

· NAG3 waste rock

- construction material for TSF in the upstream shell zone of the TSF Site D dam where practical and cost-effective

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- surplus to NAG3 dump on top of the overburden dump

- cover with compacted overburden and topsoil at closure and direct drainage towards the pit

· PAG1 and PAG2 waste rock

- subaqueous storage in TSF Site D, or

- used in downstream shell zone of TSF Site C dam below elevation 1,340 m.

· Tailings

- place in TSF Site C and D.

Figure 18-4 shows the final layout of the proposed disposal facilities.

18.4.4 Construction and Deposition Strategy

The TSF Site C main dam will be constructed to elevation 1,360 m and the TSF Site D main dam to elevation 1,345 m.  The total fill requirement for the tailings embankment construction is 78.1 Mm³, of which 5.9 Mm3 will be required for Stage 1 construction.  Construction materials will be sourced from pit stripping (75.4 Mm3 equivalent to 166 Mt) and external borrow areas (2.7 Mm3).  The embankment will be raised using the centreline method of construction once the tailing beaches have been established.  Construction of each embankment stage will be scheduled every year.

The embankment foundations will need to be cleared and stripped in preparation for fill placement for each stage.  A cut-off trench will be excavated through the surficial sands and gravels into the natural glacial till foundation to key the core zone into low-permeability foundation soils or onto bedrock.  A cut-off trench will be constructed at the TSF Site C main dam, saddle dam, and TSF Site D main dam.

Tailings from the mill process will be delivered by gravity through a pipeline from the mill to the TSF Site C for the first two years of operation.  The pipeline will be laid along a service road for start-up to the Site C main dam.  Discharge into the TSF will be directed from valved off-takes from the pipeline initially spaced at 500 m intervals.

A new delivery pipeline will be installed by the end of Year 2 to deliver tailings by gravity from the mill to TSF Site D.  The pipeline will be laid on a service road and along the main dam crest, with valved off-takes spaced at 500 m intervals.

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18.4.5 Seepage Control Measures

Seepage will largely be controlled by the low-permeability core zone constructed prior to development of the tailings beach, the tailings deposit, and the low-permeability foundation materials.  Seepage at the tailings embankment will follow the natural topography and report to seepage collection and recycle ponds at topographic low points.  The seepage recovery system for the tailings embankment will include groundwater monitoring recovery wells, seepage collection trenches, seepage collection ponds, and a seepage pumpback system.

Seepage from the TSF will result from infiltration of ponded water directly through the embankment fill and the natural ground and from expulsion of pore water as the tailings mass consolidates.  Special design provisions to minimize seepage losses include the development of extensive tailings beaches (which isolate the supernatant pond from the dam), toe drains at the downstream toe to reduce seepage gradients, and contingency measures for groundwater recovery and recycle.  An allowance for a cut-off wall has been included on the south abutment of the Site D dam.

Seepage through the embankments will be collected in the embankment filter and seepage collection systems and recycled back into the TSF.  Groundwater monitoring wells will be installed in the downstream areas as part of the monitoring program and may be converted to recovery wells if required to recover any foundation seepage.  The embankment will be constructed with toe drains to collect seepage and reduce seepage gradients.  Seepage collected in the embankment drainage systems will be transferred to seepage collection and recycle ponds at topographic low points at the downstream toe.  Additional seepage collection ditches constructed along the toe of the embankments will collect seepage and surface runoff and direct the flow to the seepage collection and recycle ponds, from where it will be pumped back to the TSF.

18.4.6 PAG Disposal Area

The PAG waste rock disposal area within the TSF footprint will be developed as part of preproduction construction to provide a location for PAG disposal from the pit stripping to expose the deposit.  The PAG disposal area will be developed at the same or similar rate of rise as the tailings but will be several metres higher to provide a dry, stable placement surface for truck traffic.  The design objective for the PAG area is to flood the waste rock within one year of placement.  The maximum elevation of the waste storage area will remain at an elevation where it can be flooded by the supernatant pond in the case of premature closure.  At closure, this waste storage area will be submerged below the final closure tailings elevation.

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The PAG disposal area will commence at the base of the TSF Site C downstream shell zone and extend outwards to the east down the valley as a large lobe, or fill platform, with overall slopes at angle of repose.  The tailings beaches will provide a low-permeability transition zone between the coarse, permeable reactive waste rock and the tailings embankments and will function as a seepage control measure

18.4.7 NAG Disposal Area

Waste dump areas have been identified and subsequently refined and optimized to minimize surface water control requirements.  NAG4 waste rock and overburden will be combined and placed in a permanent engineered stockpile adjacent to the pit on the northwest side.  Water that infiltrates through the stockpile will be collected at the base of the dump in a collection pond and directed to the TSF.  The stockpile will be developed with overall slopes to facilitate reclamation with growth media at closure.  Surface water runoff during operations and at closure will be diverted around the stockpile to the TSF.

NAG3 waste rock material not used in upstream TSF dam construction will be stockpiled on top of a sloping platform in the overburden dump to allow collection of infiltrated water for direct disposal by gravity flow into the pit.  The NAG3 stockpile will be covered with compacted overburden at the end of mine life.

18.4.8 Closure and Reclamation Considerations

Upon mine closure, surface facilities will be removed in stages and full reclamation of the TSF will be initiated.  This full closure scenario will also work well in the event of premature closure of the time.  General aspects of the closure plan include:

· Selective discharge of tailings around the facility during the final years of operations to establish a final tailing beach that will facilitate surface water management and reclamation

· Covering the exposed tailings beaches with approximately 1 m of overburden

· Pumping the tailings pond water to the pit to expedite the development of the pit lake

· Dismantling and removal of the tailings and reclaim delivery systems and all pipelines, structures, and equipment not required beyond mine closure

· Construction of overflow channels

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· Removal of the seepage collection pump-back systems at such time that suitable water quality for direct release is achieved

· Removal and re-grading of all access roads, ponds, ditches, and borrow areas not required beyond mine closure

· Long-term stabilization of all exposed erodible materials.

The seepage collection ponds and recycle pumps downstream of the dams will be retained until monitoring results indicate that any seepage from the TSF is of suitable quality for direct release to downstream waters.  The groundwater monitoring wells and all other geotechnical instrumentation will be retained for use as long-term dam safety monitoring devices.

Post-closure requirements will include an annual inspection of the TSF and ongoing evaluation of water quality, flow rates, and instrumentation records to confirm design assumptions for closure.

At closure, the waste rock dumps will be reclaimed as described above.  The low grade stockpile pad will be removed and disposed of in the open pit or TSF.  The pit will be allowed to flood and, upon filling, the pit lake will discharge via a constructed channel to the TSF and hence via constructed wetlands to Davidson Creek.

18.5 Water Management

The key facilities for the water management plan are:

· Open pit

· Mill (including fresh and process water tanks)

· Tailings storage facility (TSF)

· Diversion and water management structures

· Fresh water supply

· Sediment and erosion control measures for the facilities

The water management strategy is to utilize water within the Project area to the maximum practical extent.  The plan involves collecting and managing site runoff from disturbed areas and maximizing the recycle of process water.  Site runoff water will be stored on site within the TSF and used to inundate the potentially acid generating (PAG) tailings solids and waste rock.  Excess water will be stored in the supernatant pond within the TSF and recycled to the mill for use the process.  The water supply sources for the Project are as follows:

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· Precipitation runoff from the mine site facilities

· Water recycle from the TSF supernatant pond

· Groundwater from open pit dewatering and depressurization

· Fresh water supply from a nearby lake to supplement requirements for processing and flow reduction mitigation in lower Davidson Creek

· Fresh water from a groundwater well field for camp and potable water

· Treated black and grey water, in small quantities, from the camp.

Sediment and erosion control strategies will include establishing diversion and runoff collection ditches, constructing sediment control ponds, and stabilizing disturbed land surfaces to minimize erosion.

Activities that have the potential to require sediment and erosion control include clearing vegetation and stripping topsoil, stockpiling topsoil, and constructing roads and infrastructure.  Potential hazards from these activities, in the absence of planned mitigation measures, include increased surface erosion from disturbed areas, increased sediment loading to downstream receiving environments, and siltation or erosion of downstream watercourses or water bodies.

Sediment mobilization and erosion will be managed throughout the site by:

· Installing sediment controls prior to construction activities.

· Limiting the disturbance to the minimum practical extent.

· Reducing water velocity across the ground, particularly on exposed surfaces and in areas where water concentrates.

· Progressively rehabilitating disturbed land and constructing drainage controls to improve the stability of rehabilitated land.

· Ripping the surface in rehabilitation areas to promote infiltration.

· Protecting natural drainages and watercourses by constructing appropriate sediment control devices such as collection and diversion ditches, sediment traps, and sediment ponds.

· Restricting access to rehabilitated areas.

· Constructing surface drainage controls to intercept surface runoff.

Installation of temporary erosion and sediment control features will be the first step toward controlling sediment and erosion during construction.  All temporary sediment and erosion control features will require regular maintenance.  The temporary erosion and sediment control features will be reclaimed after the soils and sediments have stabilized.

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The construction and operations water management strategies for the Project have been developed by identifying the size and position of the planned mine site facilities, and by establishing estimated catchment area boundaries based on the proposed mine site development concept.  The approximate outlines of the proposed facilities and resulting catchment area boundaries are shown on Figure 18-5.  Runoff from all disturbed areas will drain by gravity to the TSF.

18.5.1 Sequence of Construction

Water management during construction of the TSF and associated facilities will consist of the following sequence of activities:

· Install stream bypass pumping systems at the TSF site C dam, PAG waste rock management pond dam, TSF site D dam, and environmental control dam (ECD) areas.

· Construct localized diversions to divert clean runoff around work areas.

· Install construction water infiltration ponds at each work area.

· Construct a series of upstream and downstream coffer dams at the proposed dam locations in Davidson Creek.

· Install construction dewatering systems to collect construction water and transfer it to the construction water infiltration ponds.

· Construct the ECD and install a large pump with pipeline to TSF site C dam.

· Construct the TSF site C dam.

· Construct the PAG waste rock management pond dam.

· Construct the TSF site D coffer dam.

· Install large pumps and pipelines to transfer water between storage ponds at the ECD, TSF site D coffer dam, PAG management dam, and TSF site C dam as storage capacity develops from dam construction.

All water in contact with the mine facilities, including the NAG waste dumps, open pit, and low-grade stockpile, will be collected and either conveyed to the TSF or the open pit.  The specific water management plans for the TSF, open pit, and associated facilities during operations areas are outlined below.

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18.5.2 TSF Water Management Plan

All runoff from within the TSF catchment will be stored within the TSF.

Process water will be discharged into the TSF with the tailings during operations.

A tailings deposition strategy will be implemented to selectively develop tailings beaches along the embankments, thereby producing an extensive low-permeability zone that facilitates seepage control.  The operational supernatant pond volume will be managed by selective tailings deposition to ensure that beaches are saturated, thus reducing the potential for dust generation.

The seepage collection and recycle ponds and the ECD will collect TSF seepage and sediment-laden runoff from the TSF embankments and retain the water to allow suspended sediment to settle out.  This water will then be pumped back to the TSF for use in the mill process.

18.5.3 Open Pit Water Management

Open pit dewatering will continue throughout the mine life, with dewatering flows either directed to the TSF or discharged directly to the process water tank at the mill site.

Once active mining from the pit has ceased, pit dewatering will cease and some surface water will be diverted to the open pit to contribute to pit filling.  TSF supernatant will also be pumped to the open pit, which will allow for natural degradation of contaminants in the TSF water.

18.5.4 Reclamation and Closure

All waste dump and TSF embankment benches and slopes will be reclaimed in a manner that prevents erosion and minimizes the suspension of sediments.  Water from the seepage collection ponds and ECD will continue to be pumped back to the TSF until the water quality is suitable for passive release in perpetuity.

Water diversions around the TSF will be breached to return watercourses to their natural direction at closure.  The water quality in the pit lake is expected to be acceptable for direct discharge; will be allowed to overflow to wetlands constructed in the TSF and subsequently to Davidson Creek.

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18.5.5 Water Balance

An overall average monthly site water balance assessment was carried out to determine the preliminary water management strategy and process makeup water requirements for the Project.

The results from the water balance modelling are summarized below:

· There will be no surface water discharge from the TSF to the environment during operations.

· The average annual runoff into the TSF will be approximately 10 Mm3 per year, which includes runoff from undisturbed catchments, TSF beaches, and ponds.

· Under average precipitation conditions with a start-up pond of 7 Mm3, the TSF will operate in a water surplus condition, without the need for additional makeup water, until approximately Year 4 of operations.  After Year 4, the TSF will operate in a water deficit condition and require a total of approximately 10 Mm3 of additional makeup water over the next eight years of operations.  As of Year 12, the TSF accumulates excess water for the remainder of the mine life.

· Pit runoff and pit dewatering contributes on average 1.4 Mm3 of water per year under average precipitation conditions.

· The average volume of water retained in the tailings voids, and hence not available for water recycle to the processing plant, in both TSF C and D is approximately 8 to 9 Mm3 annually.

· The volume of water retained in the PAG waste rock voids in the TSF ranges from 0.5 to 4.5 Mm3 annually.

· The tailings pond water volume for TSF C during operations will average approximately 1 Mm3 but will range between 1 and 1.5 Mm3.

· The tailings pond water volume for TSF D during operations will average approximately 7.5 Mm3 but will range between 3 and 24 Mm3.

18.5.6 Seepage Collection and Recycle Ponds

Seepage collection and recycle ponds will be constructed downstream of the TSF site C and D dams to provide collection points for seepage.  A seepage collection pond to be constructed downstream of TSF site D is designated the environmental control dam (ECD).  A temporary seepage collection pond will constructed downstream of the TSF site C dam to collect TSF seepage and PAG runoff for recycle back into the TSF site C impoundment.

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The seepage collection ponds will be constructed as water-retaining dams with appropriately designed filter and transition zones.  The ponds have been designed for the storage of seepage and surface runoff and freshet inflow.  The pump-back system will be designed for year-round operation.  HDPE pipelines laid along the downstream face and across the crest of the embankments will convey water from the seepage collection and recycle ponds into the TSF.  In the event of a pipeline rupture, these pipelines will drain back to the seepage collection and recycle ponds.

Groundwater water quality will be monitored in wells situated downstream of the embankments.  Ongoing water quality monitoring of groundwater and surface water will be used to assess the effectiveness of the seepage collection system.  The objective of the monitoring is to ensure that the environmental design criteria are being met at the monitoring points downstream of the TSF.  In the unlikely event that the seepage collection systems are found to not effectively recover seepage, then it will be necessary to install additional seepage control provisions.  The efficiency of the seepage recovery system may be improved with additional hydraulic barriers and/or interception ditches and pump-back systems.

18.5.7 Water Reclaim System

Water will be reclaimed from the tailings ponds at TSF sites C and D by barge-mounted pump stations to a land-based pump station, which in turn will deliver the reclaim water to the mill.  The water will consist of supernatant from the settled tailings and runoff from precipitation and snowmelt within the catchment area.  The floating reclaim pump stations will initially be confined in deep narrow channels at a location remote from the point of tailing discharge, maximizing the potential for recovering water of acceptable clarity.  The barge will be relocated for each stage to accommodate the rising elevation of the pond.

The reclaim pipeline from the TSF will consist of sections of large-diameter HDPE and steel pipe.  Steel pipe will be used for the initial high-pressure sections of the pipeline closest to the barge and land-based pump station, while HDPE pipe will be used for the remainder of the pipeline.

18.6 Camps and Accommodation

18.6.1 Camp Facilities

The construction camp will be sized for 800 to 1,200 personnel and will be modified and refitted into the operations camp for approximately 400 personnel.  The operations camp will have en suite bedrooms, a dining room, and recreational facilities.  Rooms will be shared during the construction phase to accommodate the larger construction workforce.  Sewage from the construction and permanent camps will be treated before disposal.

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18.6.2 Employee Housing and Transportation

Employees are expected to live in several local and regional communities.  Bus transportation to site will be provided for employees.  Further investigation based on workforce levels and logistics will be completed at a later date.  A helipad will be available on site for emergency response requirements.

18.7 Power and Electrical

Knight Piésold (KP) was retained to complete studies to determine preferred transmission line alignments that meet a balance of technical/economic viability and social/environmental impacts.  A total of six options were investigated, including five alternative 230 kV transmission line routes (Figure 18-6).  The sixth option was a 138 kV voltage line running along the preferred PEA corridor from the Glennanon (GLN) Substation.

A ranking assessment of alternatives was completed for the six transmission line options, through a collaborative effort by KP, New Gold, and AMEC E&I.  A comprehensive list of assessment criteria was developed to rank each option.  Site-specific information was not available for the options, and so the selected criteria were those that could be assessed based on readily available information.  Such information was obtained from BC Government websites Land and Resource Data Warehouse (LRDW) and iMAP BC, as well as Google Earth™.  Criteria were chosen to satisfy technical, environmental, social, schedule, and cost categories.

In addition to reviewing impacts to private land owners, the resulting five alternative corridors were reviewed by BC Hydro at a conceptual level in March 2012.  BC Hydro also provided information about the viability of the 138 kV line with interconnection at GLN Substation.  It was concluded that the risk of potential electrical instability on the 133 km long, 138 kV line was unacceptably high, and therefore this option was discarded.  BC Hydro considered the other five 230 kV transmission line routes to be viable.

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The PEA assumes construction of a 133 km long, 230 kV transmission line to supply the Project with power.  The line will have an interconnection to the BC Hydro system at the GLN Substation near Fraser Lake.  The transmission line corridor runs east from the GLN Substation for approximately 19 km and continues south for another 114 km, following existing logging roads to the Project.

While the selected PEA 230 kV GLN interconnection option ranked highest through the alternatives assessment, three of the other options had scores only marginally lower.  Small changes in individual criteria ratings or category weighting could change the outcome of this assessment, and therefore other options remain open for consideration in future studies.  Future studies will include collection of detailed environmental baseline data, completion of a BC Hydro System Impact Study, First Nations and Public consultation, and an in-depth constructability field review.  The ranking assessment will then be updated.

18.8 Fuel

Fuel will be supplied in bulk road tankers from Prince George.

18.9 Water Supply

The preliminary water balance for the Project indicates the site will experience a water deficit during operations.  Water will be required to mitigate flow reductions in Davidson Creek to sustain fisheries.  An additional water supply resource is required for the Project.

Initial consideration was given to identifying surface water drainages and groundwater sources that could be used for the Project; however, the large makeup water requirement could have adverse effects on surface water and/or groundwater sources, which in turn could affect the local fish-bearing streams.  Constructing a large water storage reservoir or a large groundwater well field was considered impractical for the site conditions, and obtaining water from larger water bodies such as lakes was determined to be more practical for supplementing makeup water during the higher freshet flow periods.  Precedence exists in British Columbia for this type of system, as it has been used successfully at the Mt. Polley Mine.

A trade-off study was carried out to examine potential water supply from lakes within close proximity of the mine site.  The best alternative was Tatelkuz Lake, approximately 15 km northeast of the Project site (refer to location on Figure 18-1).

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A 15.7 km long pipeline consisting of 18 inch diameter steel pipe is proposed.  Two stages of pumping will be required to maintain design pressure along the line because of the 416 m increase in elevation between the lake and the discharge point (TSF) at the Project site.

Tatelkuz Lake has the largest catchment and inflow, and the projected withdrawal rate would result in the least effect on flows, lake volumes, and associated environmental conditions.  Costs to install pumping and pipelines are comparable with those for the only other lake that was considered to be a viable water source

18.10 Comments on Section 18

In the opinion of the QPs, the following conclusions and comments are appropriate:

· The Project is a greenfields site and will require construction of infrastructure to support the planned producing facilities.

· Tailings and waste facility design is based on appropriate geotechnical testwork, and acid-base accounting tests.

· A number of water control structures are planned, and will include the collection and diversion of surface water runoff, sediment control ponds, and pump-back systems.

· Process and potable water will be supplied via a 15.7 km long pipeline from Tatelkuz Lake

· Road access is expected to be via the existing Kluskus Forest Service Road.

· Employees are expected to live in several local and regional communities once operations have commenced.

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19.0 MARKET STUDIES AND CONTRACTS

19.1 Market Studies

No formal marketing studies have been completed.

New Gold currently has an operative refining agreement with Johnson Matthey in Salt Lake City for refining of gold and silver doré produced from New Gold’s Mesquite Mine in Nevada.  For this operation, New Gold’s gold production is sold on the spot market, by marketing experts retained by or on behalf of New Gold.  The terms contained within the sales contracts are typical of, and consistent with, standard industry practice, and are similar to contracts for the supply of doré elsewhere in the world.

New Gold is likely to market any future production from Blackwater in the same manner as that currently employed at Mesquite.

Gold can be readily sold on numerous markets throughout the world and it is not difficult to ascertain its market price at any particular time. Since there are a large number of available gold purchasers, New Gold would not be dependent upon the sale of gold to any one customer. Gold could be sold to various gold bullion dealers or smelters on a competitive basis at spot prices.

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

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.

19.4 QP Conclusions and Comments on Section 19

In the opinion of the QPs, New Gold will be able to market gold produced from the Project. Sales contracts that could be negotiated would be expected to be within industry norms. However, the majority of production would be expected to be spot marketed.

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20.0 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

20.1 Baseline Studies

20.1.1 Air Quality and Noise

There are no significant sources of air or noise emissions in the Project area.  Studies for air quality and noise are scheduled for the latter half of 2012.

20.1.2 Terrestrial Environment

Soils and Terrain

Studies for soils and terrain are scheduled for the latter half of 2012.

Vegetation and Rare Plants

The Project footprint lies within the subboreal spruce, Englemann spruce-subalpine fir and alpine biogeoclimatic subzones.

The 2011 field work included two surveys of species at risk, a terrestrial ecosystem mapping (TEM) survey, and the collection of plant samples for baseline metal concentrations.

One Blue-listed species, whitebark pine (Pinus albicaulis), was found at high elevations in the subalpine south of the current camp location and generally outside of the proposed Project area.  Whitebark pine is listed as Endangered by the Committee on the Status of Endangered Wildlife in Canada (COSEWIC), but not under the Canadian Species at Risk Act (SARA).

No invasive plant species were found in the 2011 surveys.

Fairly extensive wetlands are present in the area, particularly at lower elevations on drainage basins originating at or near the Project site.  Water quality samples were collected in 2011 and additional studies of wetland ecological function will be completed as part of 2012 studies.  Dissolved aluminum and total cadmium arising from natural sources exceeded guidelines at a number of sites.

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Wildlife

During the 2011 surveys within and adjacent to the Project footprint, eight species of diurnal raptors were identified; nine other species of raptors were observed incidentally.  Surveys of terrestrial birds were conducted in June and July of 2011.  Diversity and abundance were higher during the June round of point counts, with 221 detections of 39 species, including two provincially and federally listed species.  Twenty-two wetland areas were surveyed, and a total of 38 individuals of nine bird species, plus one moose, were tallied.  Species diversity and population numbers encountered in 2011 are probably not representative because of the inclement weather at that time.  Additional studies are being conducted in 2012.

The Project area is east of known caribou range but includes habitat for moose, predominantly in lowlands east and north of the footprint proper.  In the survey area of Mt. Davidson, wildlife and wildlife habitat values are considered moderate, and the area does not appear to provide any critical wintering habitat for ungulates.

20.1.3 Aquatic Habitat

Surface Water Quality

Water quality has been monitored at a number of locations in the Project area (Figure 20-1.  Surface waters at the Project site have circum-neutral pH and low hardness.  Alkalinities are also low, consistent with low hardness and conductivity.  Metal concentrations are all low; typically one to several orders of magnitude below their respective BC MOE Protection of Freshwater Aquatic Life 30-day guidelines.  The few exceptions are listed below (refer to Figure 20-1 for locations):

· Total cadmium:  sites WQ1, WQ4, WQ7 exceed the maximum guideline

· Total zinc:  site WQ4 exceeds the maximum guideline

· Dissolved aluminum:  site WQ3, WQ7 exceed the 30-day guideline

· Dissolved aluminum: sites WQ1, WQ4, WQ5, WQ6, WQ10, WQ11, WQ12  exceed the maximum guideline

Water quality monitoring is ongoing.

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Groundwater Quality

Limited groundwater data are available.  The groundwater at Blackwater appears to be unusual in being soft (hardness 29 to 54 mg/CaCO3/L) and having very low alkalinity (11 to 32 mg/L), probably because of artesian conditions when sampled.  Most total metals are below BC MOE Protection of Freshwater Aquatic Life 30-day guidelines (although these guidelines do not apply), with the exceptions of cadmium, iron, and zinc (in BW101).  None of the Provincial Contaminated Sites Regulation Schedule 6 criteria that would apply off the mine site were exceeded except dissolved iron.

Stream Sediments

Stream sediments were collected once in August 2011 at water quality sampling sites.

Other than at site WQ4, metal levels were only occasionally slightly elevated above the Canadian Council of Ministries of Environment (CCME) Interim Sediment Quality Guideline (ISQG) or the BC MOE Lowest Effect Level (LEL) guidelines, with the one exception of arsenic at WQ14.  In addition, consistent with surface water quality results at WQ14, the stream sediments at this site have natural elevated levels of a number of metals, with arsenic and zinc being the highest above guidelines, exceeding permissible exposure level (PEL) and severe effects levels by 3 times (cadmium), 9 times (zinc), and 13 times (arsenic).  The WQ4 site also naturally exceeds interim sediment quality guideline/lowest effect level (ISQG/LEL) guidelines for lead and mercury.  Depending on the final mine footprint, this site may be within the contact water zone.

Fisheries

Aerial surveys of streams within the local study area (LSA) were initially conducted in spring 2011 to identify fish barriers and validate reach breaks.  Established reach breaks were also verified in the field during the summer field program.  Stream habitat assessments and a hoop net program were conducted for creeks in the Project area watersheds.

No fish barriers were identified during spring aerial surveys and the summer field program.  Based on knowledge from baseline field programs in 2011 and early 2012, fish habitat is widely distributed, with fish access available into headwater drainages within the Project footprint and affected drainages.

Rainbow trout distribution was found to be widespread throughout all tributaries sampled in 2011.  With the exception of one small creek, rainbow trout were captured in all reaches of streams sampled by backpack electrofishing in summer 2011.  Lake surveys showed that all lakes surveyed support populations of rainbow trout and longnose sucker; a number of smaller ponds remain to be sampled.

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Kokanee salmon spawning runs were documented during late August 2011 in lower Davidson Creek and Creek 661.

20.1.4 Archaeology

An archaeological overview assessment (AOA) has been completed in the Project area.  Culturally modified trees were identified in one location near the Project area.  No other archaeological sites, historic heritage sites, or evidence of any previous archaeological investigations were identified within the Project area.

The regional study area (RSA) consists of a 10 km buffer around the Project area.  Within this area, 193 archaeological sites have been identified.  General trends suggest that these sites are situated at lower-elevation settings in proximity to aquatic features, pine stands, larger water bodies, and / or in association with the Blackwater River drainage (which will not be affected by the mine development).

An Archaeological Impact Assessment as per provincial guidelines will be conducted in during 2012–2013.

20.2 Environmental Issues

The approach for designing the Project incorporates environmental principles intended to avoid, minimize, mitigate, and compensate (when the avoidance, minimization, or mitigation are not sufficient) the impacts of the Project on the environment.  To this end, the following design principles will be applied to optimize the environmental performance of the Project and to achieve the requirements for environmental protection with the demands of construction and operation:

· Massing

· Clustering

· Incorporating Traditional Knowledge and land use

· Identifying key environmental constraints and opportunities

· Preventing surface water discharge during operations

· Designing for closure

· Ongoing risk assessment and decision-making.

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“Massing” consists of the early identification and avoidance of key sensitive areas in the region where the Project is located.  These include:

· The Blackwater River drainage, which will be avoided by any project facility;

· The Ungulate Winter Range, which is the winter habitat for the Tweedsmuir-Entiako Caribou herd and will be avoided by any project facility; and

· Primary growth areas for the whitebark pine tree, which is a Blue-listed species.

“Clustering” of the Project facilities minimizes the overall footprint of the Project and ensures that the mine waste management facilities such as the tailings storage facility and the waste rock dumps are located in proximity to the open pits and within the same or adjacent sub-drainages.  Use of the existing access roads also minimizes the overall footprint of the Project and has been included in the Project design.

Traditional Knowledge and Land Use (TK/TLU) studies provide information about the value of the different environmental and cultural components to First Nations.  This information is being collected by engaging local First Nations to complete a comprehensive TK/TLU study.  These studies are expected to be completed in late 2012 to early 2013.  Archaeological studies will also provide information about the existence of areas of high archaeological potential to be avoided, where possible, by the Project facilities.  In addition, existing recreational and tourism use areas have been identified in proximity to the Project, including the Alexander Mackenzie Heritage Trail, Messue Wagon Road, and Messue Horse Trail / Kluskus Bypass. These areas will be avoided by any Project facility.

A key environmental consideration for the Project is the presence of fish-bearing habitat.  Fisheries values are higher in the lower reaches of Davidson Creek, where Kokanee salmon spawning habitat has been identified.  The Project area avoids these lower reaches.

The Project will prevent surface water discharge during the operations phase.  Measures include diverting water to the tailings storage facility (TSF) pond and collecting potential seepage from mine water facilities.  Also, process water will be recycled to minimize the requirements for fresh water makeup.

Designing for closure is a principle that will ensure the long-term potential impacts of the Project are avoided, minimized, and mitigated appropriately.  This involves early consultation with stakeholders and First Nations regarding closure objectives as well as applying mine waste management criteria that involve segregation, encapsulation, or underwater disposal to minimize the potential for long-term impacts.  In consideration of this principle, the waste rock dumps and open pit will be designed to drain towards the TSF, and passive treatment options, such as constructed wetlands, will be incorporated.

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The potential environmental risks and opportunities will be assessed to keep the risks under acceptable levels and to maximize opportunities for positive outcomes.

A formal environmental impact assessment will be completed for the proposed Project and a comprehensive assessment of the potential environmental effects of the Project will be conducted.  This section addresses the key potential issues, but only at an overview level.

The potential environmental effects are associated with potential impacts to environmental receptors from construction, operation, and closure of the proposed Project and include:

· Loss of fish habitat in Davidson Creek and tributaries that will need to be mitigated where possible and compensated for where necessary under the Canadian federal Fisheries Act in accordance with Fisheries and Oceans No Net Loss Policy.

· Permanent changes in natural landforms where the open pit, waste rock pile(s), tailings impoundment, and gravel borrow source are located.

· Potential change in water quality and quantity in Davidson Creek due to construction and operation of the tailings impoundment.

· Potential effects of open pit development on the white bark pine ecosystem at high elevations of Mt. Davidson.

· Temporary disturbance of other areas, which will be rehabilitated to pre-mining landforms on reclamation and closure of the mine.

· Environmental monitoring during construction, operation, and closure.

20.3 Environmental Liabilities

20.3.1 Existing Environmental Liabilities

Existing environmental liabilities are related to the exploration activities that New Gold has undertaken to date or to activities conducted by third parties prior to New Gold’s acquisition of the Blackwater mineral claims.  A Phase 1 Environmental Site Assessment was conducted for the proposed mine site, and no records of site activities were identified prior to 1973.  It is confirmed that the potential for environmental liabilities is limited to exploration activities conducted since 1973.  These potential liabilities are within the reclamation bond amount.

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20.3.2 Exploration Reclamation Bonding

A reclamation plan for the exploration activities was prepared to accompany the Notice of Work and Reclamation application to the BC Ministry of Energy and Mines.  The main reclamation objective is to return the site to a state similar to its pre-disturbance appearance and functionality.  The reclamation plan addresses drilling, geophysical surveys, access to site, and the camp and camp services, and proposes a series of measures that include removing buildings, machinery, debris, or contaminated soils; re-establishing natural landscapes; and deactivating roads and trails.  The current established security deposit for project reclamation is approximately $1 million.

20.3.3 Other Environmental Liabilities

New Gold has signed a Road Maintenance Agreement with Canfor Forest Products Ltd. for the use of 150 km of the Kluskus Forest Service Road to access the Project, but it is only responsible for payment of a portion of the maintenance costs.  No existing environmental liabilities along the road are the responsibility of New Gold.

The potential for other environmental liabilities (not related to exploration activities) is relatively low, due to the remote location of the Project and that mineral exploration in the Blackwater mineral claims started in recent years and there is no evidence of previous mining activities.

20.4 Closure Plan

Knight Piésold has estimated a mine closure cost of $101 million.  This cost includes reclamation and closures of the TSF and waste dumps, decontamination of facilities as required, removal of all constructed surface facilities, removal of the power and pipelines, closure and reclamation of access roads, and post-closure monitoring and maintenance to ensure that end land use objectives are met.

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20.5 Permitting

The Project is subject to review under the BC Environmental Assessment Act (British Columbia) and the Canadian Environmental Assessment Act (Canada).  To initiate the Environmental Assessment (EA) process a project description has been written in parallel with completion of the PEA, and is due for submission late 2012.  Draft Terms of Reference for the provincial Environmental Assessment Certificate Application (EAC) draft guidelines for the federal EIS are expected to be filed in Q1 2013 for approval by the BC Environmental Assessment Office (BCEAO) and CEAA.  An EAC Application/EIS will then be filed, and reviewed with the applicable agencies, First Nations and stakeholders.  Following completion of this review, New Gold anticipates receiving required EA approval in mid to late 2014.

Following EA approval, additional permits, licences, authorizations, and certificates will be required to proceed to construction of the Project.  New Gold has identified and commenced planning for obtaining all such approvals in parallel with the EA approval process.  These latter approvals are expected by early 2015.  Table 20-1 and Table 20-2 present the preliminary list of required permits, licences, authorizations, and certificates under provincial and federal regulations.

Table 20-1: BC Provincial Permits, Licences, and Authorizations

Statute	Authorization or Requirement	Agency
BCEAA	EA Certificate	BC EAO
Condition of provincial EA	Fish and Wildlife Mitigation and Monitoring Plans	BC EAO or BC MFLRNO
Mines Act  and  Environmental Management Act	Joint Application for a Mines Act Permit and Effluent Discharge Permit	BC MEM and BC MOE
Forest Act	Occupant Licence to Cut - Sec 47	BC MFLNRO
Forest Protection Code (FPC) Act, Provincial Forest Use Regulations, Forest and Range Practice Act	Special Use Permit	BC MFLNRO
FPC Act, Provincial Forest Use Regulations, Forest and Range Practice Act	Road Use Permit	BC MFLNRO
Water Act, Water Regulation	Approval or Notification of “changes in or about a stream” (s.8/s.9)	BC MFLNRO
Water Act, Water Protection Act	Water Licence	BC MFLNRO
Heritage Conservation Act	s. 14 Inspection Permit	BC MFLNRO
Heritage Conservation Act	s. 14 Investigative Permit	BC MFLNRO
Heritage Conservation Act	s. 12 Site Alteration Permit	BC MFLNRO
Environmental Management Act	Fuel Storage Permit	BC MOE

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Statute	Authorization or Requirement	Agency
Environmental Management Act - Hazardous Waste Regulation	Hazardous Waste Registration	BC MOE
Environmental Management Act	Air Discharge Permit	BC MOE
Environmental Management Act	Refuse Permit	BC MOE
Environmental Management Act - Municipal Sewage Regulation	Sewage Registration	BC MOE
Wildfire Act	Burning Permit	BC MFLNRO
Wildfire Act	Amendment to Closed Area Regulations	BC MFLNRO
Transportation Act, Motor Vehicles Act	Access Permit (MOTI-A)	BC MOTI
Motor Vehicles Act	Approvals for oversize loads or bulk haul	BC MOTI
Drinking Water Protection Act and Regulation	Construction Permit - Sec 2	NHA
Drinking Water Protection Act and Regulation	Operating Permit - Part 2	NHA
Public Health Act – Food Premises Regulation, and Drinking Water Protection Act	Food Premises Permit	NHA
Public Health Act – Sewage Disposal Regulation	Filing of Certification Letter - Sec 32	NHA
Safety Standards Act	Permit	BC Safety Authority
Wildlife Act	Fish or Wildlife Permits	BC MFLNRO
Mining Right of Way Act	Mining Right of Way Permit	BC MEM
Land Act	Investigative Use Permit	BC MFLNRO
Land Act	Licence of Occupation	BC MFLNRO
Mines Act	Explosives Storage and Use Permit	BC MEM

Note: BCEAA - British Columbia Environmental Assessment Act; BC EAO - British Columbia Environmental Assessment Office; BC MEM - British Columbia Ministry of Energy and Mines; BC MFLRNO - British Columbia Ministry of Forests, Lands and Natural Resource Operations; BC MOE - British Columbia Ministry of Environment; BC MOTI - British Columbia Ministry of Transportation and Infrastructure; EA - Environmental Assessment; EHO - Environmental Health Officer; EP - Environmental Protection; FPC - Forest Practices Code; HCA - Heritage Conservation Act; n/a - not applicable; NHA - Northern Health Authority; ROW - Right-of-Way

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Table 20-2: Federal Permits, Licences, and Authorizations

Statute	Authorization or Requirement	Agency
CEA Act, 2012	EA Decision - Comprehensive Study (EA-Canada)	CEA Agency
MMER under Fisheries Act	Compliance and Reporting	EC
Fisheries Act	Authorizations under s.35(2) - Approval of final FHCP (HADD) Authorization under s.36(3) – Schedule 2 Amendment under the MMER	DFO
Explosives Act; Explosives Regulations and Ammonium Nitrate and Fuel Oil Order Regulation	Licences under s.7(1)(a) Mechanical Ammonium Nitrate & Fuel Oil Certificate Assemble and Blend Ammonium Nitrate & Fuel Oil	NRCan
Radio Communication Act	Licenses	Industry Canada
Transportation of Dangerous Goods Act; Transportation of Dangerous Goods Regulations	Transportation of Dangerous Goods Permit	Transport Canada
Navigable Waters Protection Act	Transport Canada’s approval	Transport Canada

Note:  CEA Agency - Canadian Environmental Assessment Agency; CEA Act, 2012 - Canadian Environmental Assessment Act; 2012; DFO - Fisheries and Oceans Canada; EA - Environmental Assessment; EC - Environment Canada; FHCP - Fisheries Habitat Compensation Plan; HADD - harmful alteration, disruption or destruction; MMER - Metal Mining Effluent Regulations; n/a - not applicable; NRCan - Natural Resources Canada

20.6 Considerations of Social and Community Impacts

20.7 First Nations Engagement

The land area involved in the Project (like much of British Columbia) is an “unsettled area” – meaning treaties were not signed historically between the Crown (the federal and/or provincial governments) and First Nations.1  In the absence of treaties, the precise nature of the Aboriginal rights held and the precise location of Aboriginal title remains undefined.  These Aboriginal groups have asserted Aboriginal rights and title over large tracts of Crown land.  Many of these asserted “traditional territories” overlap with neighbouring claims.

Pending settlement of claims to Aboriginal rights and title, the Crown (federal and/or provincial governments) has a legal duty to consult with and, if necessary, accommodate Aboriginal groups that arises “when the Crown has knowledge, real or constructive of the potential existence of the Aboriginal right or title and contemplates conduct that might adversely affect it.”2

 

1 This is in contrast to much of the rest of Canada, where treaties were signed (1871 to 1923) between First Nations and governments.  These historic treaties established Indian Reserves and provided that Aboriginal rights and title to the remainder of (Crown) land were extinguished.  In British Columbia, Indian Reserves were created, notwithstanding the absence of treaties.  Even though most of the land mass of British Columbia is not covered by treaties, there are treaties in respect of the north eastern corner (which is covered by Treaty Number 8), the Douglas treaties on southern Vancouver Island, and modern treaties concluded with the Nisga’a, Tsawwassen, Maa-nulth and Yale First Nations.

 

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The intent of engagement with First Nations is to increase the mutual awareness and understanding of the Project and its potential effects and to explore potential future strategies to mitigate negative effects and enhance positive ones.  To that end, New Gold has engaged in a systematic process to identify, contact, and communicate with First Nations.

Engagement with First Nations will continue to be an important aspect of the Blackwater project.

20.8 Other Stakeholder Consultation

Stakeholders include Federal, Provincial, and local government agencies, and land and resource users.  The nearest residence is more than 10 km away from the Project area.  The intent of the various consultations is to increase the mutual awareness and understanding of the Project and its potential effects and to explore future potential strategies to mitigate negative effects and enhance positive ones.  To that end, New Gold has engaged in a systematic process to identify, contact, and dialogue with these parties.  New Gold is engaged in ongoing consultation with Provincial and Federal ministries and agencies about research design, permits, and environmental assessment questions and issues as they arise.

Consultation with governments, local municipalities, forestry companies, and other land users in the area and will continue as Project planning advances.  New Gold remains committed to undertaking this consultation to the highest standards, increasing both the scale and scope of interactions in the short to intermediate term.

20.9 Discussion on Risks to Mineral Resources and Mineral Reserves

The Project is subject to Federal and Provincial permitting requirements, which include public and First Nations consultations.  The BCEAA and CEAA, 2012 have specific, regulated timelines, and in addition, provincial and federal permits have specific timelines and requirements that must be met.  New Gold has outlined a permitting schedule based on these timelines; however changes to the schedule may occur as a result of the consultation processes and any subsequent amendments that may be then required to support furtherance of permit application processes.

 

2 Haida Nation v. British Columbia (Minister of Forests), [2004] 3 S.C.R. 511 at para. 35.

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20.10 Comments on Section 20

In the opinion of the QPs, the following conclusions and comments are appropriate:

· Baseline studies have been initiated and are ongoing.

· Dissolved aluminum and total cadmium arising from natural sources exceeded guidelines at a number of sites.

· Culturally modified trees were identified in one location near the Project area.  No other archaeological sites, historic heritage sites, or evidence of any previous archaeological investigations were identified within the Project area.

· Mine closure costs based on the assumptions in this Report are estimated at $101 million

· The Project is subject to review under the Environmental Assessment Act (British Columbia) and likely the Canadian Environmental Assessment Act (Canada).  To initiate the EA process a project description has been written in parallel with completion of the PEA and is due for submission in late 2012.  Following EA approval, additional permits, licences, authorizations, and certificates will be required to proceed to construction of the Project.  New Gold has identified and commenced planning for obtaining all such approvals in parallel with the EA approval process.

· Existing environmental liabilities are related to the exploration activities that New Gold has undertaken to date or to activities conducted by third parties prior to New Gold’s acquisition of the Blackwater mineral claims

· New Gold has made contact with First Nations groups who have interests in the Project area.

· New Gold is engaged in ongoing consultation with provincial and federal ministries and agencies about research design, permits, and environmental assessment questions and issues as they arise

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21.0 CAPITAL AND OPERATING COSTS

21.1 Initial Capital Cost Estimates

21.1.1 Basis of Estimate

The capital cost estimate for the Project covers the costs to design, procure, construct, and commission the facilities described in this Report.  The estimate is categorized as an AMEC Class 5, Scoping Level Estimate in Q2 2012 Canadian dollars, with an expected accuracy of ±25%.  AMEC Class 5 studies typically have an estimate accuracy range of ±25% to 50%.  The estimate is in line with the Association for the Advancement of Cost Engineering (AACE) guidelines for Concept Screening Estimates, which have a lower range of accuracy recommendation of -20% to -50% and an upper range of accuracy recommendation of +30% to +100%.

The estimate reflects the combined efforts of New Gold, AMEC, Independent Mining Consultants Inc. (IMC), Knight Piésold Consulting Inc. (KP), and Allnorth Consultants Limited (Allnorth).  Each consultant provided input to the capital cost estimate, appropriate to a Class 5 estimate.  Areas of responsibility are indicated in Table 21-1.  AMEC was responsible for coordinating the data and assembling the estimate.

 Table 21-1:  Capital Cost Estimate Responsibility

Consultant	Responsibility
AMEC	Designs and estimates for site layout, surface materials handling, process plant, ancillary buildings, on-site utilities, power distribution, control systems, spares, over-all compilation of all direct costs, construction schedule, construction indirect costs, contingency.
Independent Mining Consultants (IMC)	Mining preproduction and mining equipment capital cost
Knight Piésold (KP)	Water supply, tailings management facilities, off-site transmission line.
Allnorth	Access road cost.
New Gold	Owner’s cost.

Capital costs for surface facilities include the purchase of materials and equipment, construction and installation of all structures, utilities, materials, and equipment, and all associated indirect and management costs.  They also include contractor and engineering support to commission the process plant to ensure all systems are operational.  At the point of handover of the plant to the Owner’s operations group, all subsequent costs, including ramp-up to full production, are considered operating costs. 

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The capital cost estimate was based on the following:

· Process flow diagrams

· Project scope of facilities

· Process plant equipment list

· Design criteria

· Conceptual site layout drawings

· Conceptual general arrangement drawings

· Electrical single-line diagrams

· Supplemental sketches where required;

· Geotechnical and hydrogeological reports

· Regional climactic data

· Project work breakdown structure (WBS)

· AMEC’s In-house data.

No allowance was included in the estimate for escalation beyond second quarter 2012.

The estimate was based on a combination of semi-detail estimating for some components of the Project and capacity factor or equipment factor estimating for others.  The estimate includes:

· Direct field cost of executing the Project; and

· Indirect cost associated with the design, construction, and commissioning of the facilities.

The estimate excludes off-site facilities, such as an off-site reagent transfer and storage facility.

An order-of-magnitude estimate of overall on-site man-hours is approximately 6 million hours.  A detailed breakdown of the man-hours is not provided at this stage of study.

Contingency is included in the estimate as a percentage of direct and indirect costs.  Separate contingency amounts are included for mining, tailings storage facility, off-site power line, and all other direct costs, as there were different levels of certainty and estimation applicable for different areas of the scope.  A deterministic method was used to calculate contingency for AMEC scope.  The level of contingency on AMEC scope is slightly lower than normally assigned to a scoping study estimate, due in part to AMEC’s ability to use in-house process equipment costing data from relevant benchmark projects in B.C.  No contingency is applied to Owner’s costs.

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Currency conversion rates were developed by New Gold, and assumed 1 CAD = 1.00 USD, and 1 CAD = 0.714 Euro.

21.1.2 Mine Capital Costs

Mine design, mine scheduling, and mine fleet sizing and selection were performed by IMC.  Estimated initial mine capital costs total $208 million (Table 21-2).

 Table 21-2:  Mine Capital Costs

Area	Description	Cost ($M)
1111	Mine Preproduction Development	55.3
1132	Open Pit Mine Equipment	146.5
1133	Mine Dewatering and Drainage Equipment	2.0
1135	Explosives Handling and Storage (Site Preparation)	0.6
1136	Mine Pit Electrification	4.0
Total 1000	Mining	208.4

Capital mining equipment includes all drill, shovel, loader, and truck fleets, dozers, graders, and other major and minor support equipment for the mine.  The fleet requirements were calculated based on the mining movement schedule and nominated waste dump design, which determines required haulage distances.  The initial capital equipment is that required for the preproduction mining and 40% of the capital cost for equipment needed to complete Year 1 material movement.  The provision for Year 1 equipment allows sufficient funds to place deposits on selected equipment and commence assembly on site in time for Year 1 production requirements.

Provisions are included for the explosives plant site preparation based on in-house data, and mine dewatering based on Knight Piésold in-house data.

21.1.3 Infrastructure Capital Costs

On-site infrastructure includes the following areas:

· General site preparation and drainage,

· On-site access roads,

· Plant mobile equipment,

· Fuel storage area preparation,

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· Fresh, potable, process, gland, and fire water supply and distribution,

· Electrical substation, distribution and control systems, and emergency power supply,

· Permanent accommodation complex.

· Ancillary buildings and facilities: assay laboratory, administration, change-house, cold storage, truck shop/warehouse, truck wash, lighting, solid waste disposal, sewage collection and treatment, medical facilities, and security.

Costs are summarized in Table 21-3.

 Table 21-3:  On-Site Infrastructure Capital Costs

Area	Description	Cost ($M)
2110	Site Preparation & Drainage	31.0
2120	On Site Access Roads	5.4
2130	Plant Mobile Equipment	4.1
2142	Fuel Storage and Distribution System	1.9
2151	Fresh Water System and Water Supply Wells	4.7
2152	Potable Water System	2.0
2153	Process Water Storage and Distribution System	1.2
2154	Fire Water System	2.9
2155	Gland Seal Water	0.1
2156	Sewage Collection and Treatment System	4.4
2157	Solid Waste Disposal	4.8
2161	Distribution Control System	7.0
2183	Main Sub Station	19.8
2184	Power Generation	4.4
2191	Surface Power Distribution	4.4
2192	Yard Lighting	0.2
2220	Administration Building	5.7
2225	Truckshop	19.0
2240	Warehouse Building	7.2
2250	Assay Laboratory	5.0
2260	Security Gatehouse and Weigh Scale	0.7
2270	Medical Facilities	3.0
2271	Cold Storage	2.1
2272	Accommodation Complex	39.6
Total 2000	On-Site Infrastructure	180.8

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21.1.4 Process Capital Costs

Capital costs for the process plant include facilities for primary crushing, coarse ore conveying and stockpiling, grinding, whole ore leaching, gold recovery circuits, reagents, and the process building.  These costs are broken down in Table 21-4.  Table 21-5 shows the installed cost for major equipment items in the process plant.

 Table 21-4:  Process Plant Capital Costs

Area	Description	Cost ($M)
3110	Primary Crushing	34.7
3120	Coarse Ore Conveying	21.6
3130	Coarse Ore Stockpile and Reclaim	33.3
3210	Grinding Circuit	139.2
3212	Pebble Crushing	25.3
3214	Gravity Concentration	5.1
3270	Reagent Storage and Mixing	6.2
3293	Plant Air System	1.0
3296	Construction Equipment	39.6
3299	Process Building	83.5
3415	Whole Ore Leaching	118.6
3445	Carbon and Gold Recovery Circuit	26.3
3480	Cyanide Destruction Flotation and Dewatering	4.8
Total 3000	Process Plant	539.1

 Table 21-5:  Major Process Plant Equipment Costs

Area	Equipment	Number of Units	Cost ($M)
Crushing	Primary Gyratory Crusher – 60” x 89”	1	6.6
Grinding	SAG Mill – 36’ diameter	2	41.4
	Ball Mill – 24’ diameter	2	32.0
Pebble Crushing	Cone Crusher – XL1300	2	6.9
Whole Ore Leaching	Leach Feed  Thickener – 80 m diameter	1	7.0
	Residue Thickener – 80 m diameter	1	7.0

21.1.5 Waste Management Costs

Capital costs for the waste management facility were estimated by Knight Piésold and included the following:

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· Site preparation – logging, clearing and grubbing, topsoil stripping, access road development, construction dewatering, and sediment and erosion control using best management practices (BMPs),

· Tailings facility – dam footprint foundation preparation, foundation and embankment drains, embankment fill, seepage recovery systems, and basin foundation preparation,

· Tailings pipeline from process plant to TSF discharge,

· Seepage collection ponds,

· Electrical – power line and electrical equipment,

· Mechanical – pump stations, reclaim barge, tailings, and reclaim piping for reclaiming water for recycle to the process plant,

· Geotechnical instrumentation.

Where possible, the mine truck fleet will be used to place material onto the TSF embankments.

The breakdown of capital costs for waste management is shown in Table 21-6.

 Table 21-6:  Waste Management Capital Costs

Area	Description	Cost ($M)
4110	Tailings Disposal	51.0
4140	Water Management	14.4
4150	Low Grade Stockpile and NAG3 Disposal	4.8
4160	NAG4 / Overburden Disposal	0.9
4170	TSF / Waste Mobilization/Demobilization	2.8
Total 4000	Waste Management	73.8

21.1.6 Road Access Capital Costs

Capital cost estimates were developed for upgrading existing roads and constructing new sections of road to access the Project site.  It was estimated that 24.2 km of the existing Kluskus Forest Service Road would need to be widened from 6 m to 9 m and re-surfaced.

A new 24.2 km road section will be required from the Kluskus FSR to the Project site.  Costs for the new road section include the upgrade of one bridge crossing.  New road construction costs assume access to suitable gravel borrow pit material within a 10 km distance from the road construction corridor.

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The direct capital cost for the access road upgrades and construction of a new 24.2 km long road from Kluskus FSR to the Project site is estimated at $15.7 million.

Off-Site Infrastructure Capital Costs

Off-site infrastructure, estimated by Knight Piésold, comprises the power interconnection and transmission line, and the water supply infrastructure from Tatelkuz Lake to the Project site.

The preferred PEA power line option has an interconnection at the BC Hydro Glennanen Substation and a 133 km, 230 kV power transmission line to the Project site.  BC Hydro estimated the interconnection costs, including a substation upgrade, breaker, and switchgear, at $7 million.  The transmission line was estimated at $42.6 million.  This includes:

· Right-of-way clearing

· Survey of structure locations

· Structures supply and erection (poles, insulators, and all other hardware)

· Conductor supply and stringing

· Construction and maintenance access roads

· Foundations

· Grounding and shield wire where required

· Mobilization/demobilization

· Commissioning.

The capital costs of water supply from Tatelkuz Lake were estimated at a preliminary order-of-magnitude level, to supply up to 6 Mm3/a of water to the Project site.  The pipeline is assumed to be 18” diameter standard carbon steel and forms the majority of the capital cost.  Costs are summarized in Table 21-7.

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 Table 21-7:  Water Supply Capital Cost Breakdown

Item	Unit Cost	Quantity	Cost ($M)
Supply and Install 18’ dia Steel Pipe	$650/m	15,700 m	10.2
Logging	$2,050/ha	2.4 ha	0.1
Pioneer Road	$100,000/km	2.4 km	0.2
Power Lines	$100,000/km	15.7 km	1.6
Lake Intake Structure	$1,000,000 each	1	1.0
Pump Stations	$3,000,000 each	2	6.0
Construction Mobilization/Demobilization			0.8
Total			19.8

21.1.7 Owner’s Capital Costs

Owner’s cost is estimated at 5% of direct capital cost, for a total of $54.3 million.

21.1.8 Indirect Capital Costs

Indirect costs include:

· Engineering Procurement and Construction Management (EPCM)

· Temporary construction facilities and services

· Construction equipment (cranes over 90 tonnes)

· Construction camp rental and catering services

· Freight and logistics

· Capital spares, commissioning spares, first fills of reagents and grinding media, vendor representatives, and contractor support during commissioning.

Table 21-8 shows the breakdown of indirect costs and the respective percentages of direct capital costs, excluding mining.  Where applicable, different EPCM ratios were applied for the different areas of scope (excluding mining), as shown in Table 21-9.

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Table 21-8: Indirect Costs

Area	Description	% Direct Costs excl. Mining	Cost ($M)
9210	EPCM	12.7	112.0
9220	Temporary Construction Facilities	3.5	31.0
9230	Temporary Construction Services	5.5	48.7
9240	Construction Equipment	1.3	11.5
9260	Construction Camp & Catering	5.7	54.6
9270	Freight / Logistics / Taxes / Duties	4.7	44.2
9281	Spares / Warehouse Inventory	1.2	10.7
9285	First Fills	0.5	4.6
9286	Vendor Reps	0.5	4.9
9288	Pre-Commissioning and Commissioning	0.5	4.6
Total 9200	Indirect Costs	36.3	326.7

 Table 21-9:  EPCM Costs by Area

Area	% Directs	($M)
Knight Piésold – TSF and Waste	12.5	9.2
Knight Piésold – Transmission Line	8.0	4.8
Knight Piésold – Water Supply	13.0	2.8
Allnorth – Access Roads	7.0	1.1
AMEC – Other excl. mining	13.0	94.1
Total EPCM		112.0

21.1.9 Contingency

Contingency is included in the estimate as a percentage of direct and indirect costs.  Considering the different levels of certainty and estimation applicable for different areas of the scope, separate contingency amounts are included for mining, the tailings storage facility, the off-site power line, and all other direct costs.  No contingency is applied to Owner’s costs.  The total contingency for the overall estimate is $345 million, or an average of 23.5%.

Table 21-10 shows a breakdown of the contingency estimate and the percentage applied to respective sections of the estimate.

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 Table 21-10:  Contingency Costs

Area	Description	% Contingency	Cost ($M)
1000	Mining	15.0	30.3
2000 - 3000	On-site Infrastructure and Processing Plant	25.0	180.8
4000	Waste Management (TSF)	35.0	29.1
5000	Project Access Corridor	25.0	4.2
7000	Off-site Infrastructure	30.0	23.7
9100	Owner’s Costs	0.0	0.0
9200	Indirect Costs	25.0	77.6
Total PP	Contingency	23.5	345.6

21.2 Sustaining Capital

IMC developed a schedule for sustaining capital for mining equipment over the operating life of mine.  A provision for sustaining capital cost for the process plant, site infrastructure, and TSF has been determined.  The sustaining capital schedule by year is shown in Table 21-11 and is included in the economic model.  The estimated LOM sustaining capital expenditure is $571 million.

 Table 21-11:  Sustaining Capital Schedule

Operating Year	Mining ($M)	TSF/Waste ($M)	Other ($M)
1	46.5	62.4	0.0
2	4.9	29.2	5.0
3	98.3	30.5	5.0
4	1.6	13.4	5.0
5	0.1	9.1	5.0
6	21.2	9.7	5.0
7	17.8	9.4	5.0
8	2.4	7.8	5.0
9	0.2	6.2	5.0
10	2.1	6.9	5.0
11	0.0	5.1	5.0
12	48.2	6.6	5.0
13	37.1	4.4	5.0
14	8.4	4.5	5.0
15	0.0	4.6	0.0
16	0.0	5.6	0.0
17	0.0	2.2	0.0
Total Life of Mine	288.8	217.7	65.0

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

21.3 Mine Closure Salvage Value

It is assumed that $77 million salvage value can be obtained from major mine equipment and the crushing and grinding equipment from the processing plant.  Table 21-12 shows the breakdown of expected salvage value.

 Table 21-12:  Salvage Value

Unit	Purchase Cost New ($M)	Salvage %	Salvage Value ($M)
Shovels	76.2	25%	19.1
Loader	4.3	15%	0.6
Haul trucks	119.4	20%	23.9
Drills	13.6	15%	2.0
Dozers	4.8	15%	0.7
Primary crusher	6.6	35%	2.3
SAG Mills	41.4	35%	14.5
Ball Mills	32.0	35%	11.2
Pebble Crushers	6.9	35%	2.4
Total Salvage Value			76.8

21.4 Capital Cost Summary

The total estimated cost to design, construct, and commission the 60,000 t/d gold facilities described for the Project is $1,814 million, and is summarized in Table 21-13.  A parity USD:CAD exchange rate has been assumed for the purpose of determining the capital cost in US$, therefore an initial capital cost of US$1,814 million is assumed.

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 Table 21-13:  Initial Capital Cost Summary by Major Area

Area	Description	Cost ($M)
1000	Mining	208
2000	On-Site Infrastructure	181
3000	Processing Plant	539
4000	Water/Waste Management	74
5000	Project Access Corridor	16
7000	Off-Site Facilities	69
	Subtotal Direct Cost	1,087
9000	Indirect Cost	327
9100	Owner’s Cost	54
	Subtotal Indirect & Owner’s Costs	381
	Total Direct, Indirect & Owner’s Costs	1,468
PP00	Contingency	346
	Total Project Cost	1,814

21.5 Operating Cost Estimates

21.5.1 Basis of Estimate

A total labour loading factor of 32% has been applied to labour costs to allow for benefits including vacation, retirement savings, pension plan contributions, unemployment insurance, workers compensation, sick leave, and training.

Consistent energy cost assumptions have been applied to mining, processing, and general and administration costs.

A diesel cost of $1.04/L has been assumed, based on an average oil price of US$90/bbl and exchange rate of CAD $1 = US$ 0.94.

An average electricity tariff rate of 4.44 c/kWh was assumed for the Project, based on the BC Hydro tariff rate effective at May 2011, Schedule 1823: Stepped Rate, BC Zone 1 (BC Hydro, 2012).

Power consumption has been estimated for the mining, processing, and G & A areas, and power costs have been allocated proportionally.

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21.5.2 Mine Operating Costs

Mining operating costs were estimated by IMC and include the following items:

· Drilling

· Blasting

· Loading

· Hauling

· Auxiliary

· General mine

· General maintenance

· Administration for the Mining Department.

The mine will operate on two 12 hour shifts per day, seven days a week for 355 scheduled days per year, with an allowance of 10 lost days per year for weather and other impacts.  The mine operations and maintenance departments will have a salaried and supervisory staff of 40 people, including:

· Mine Manager and Secretary (2)

· Mine Operations (10)

· Mine Maintenance (10)

· Mine Engineering (8)

· Mine Geology and Ore Control (10).

The hourly personnel head count begins during preproduction with 89 in operations and 59 in maintenance.  The number of hourly persons peaks in Year 9 with 237 in operations and 146 in maintenance.

Different unit costs have been estimated for these locations, as summarized in Table 21-14, and include allocations for all the components of costs listed above.

The drills and cable shovels will be electric with the maximum use of 28.6 million kWh in Year 5; all of the remaining mine equipment will use diesel fuel with its peak consumption at 45.6 million litres in Year 7.  The LOM operating costs can be split into the major categories as shown in Table 21-15.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

 Table 21-14:  Mining Unit Costs for Different Material Types and Dump Locations

Material Type	Destination	Operating Unit Cost ($/t material)
Mineralization	Open pit to mill	1.64
	Open pit to low-grade stockpile	1.42
	Low-grade stockpile to mill	1.16
Overburden	Pit to TSF embankment	2.09
	Pit to waste dump	1.40
NAG 4	Pit to TSF embankment	2.19
	Pit to waste dump	1.71
NAG 3	Pit to TSF embankment	2.17
	Pit to waste dump	1.78
PAG 2	Pit to TSF embankment C	1.83
	Pit to TSF pond	2.00
PAG 1	Pit to TSF embankment C	1.82
	Pit to TSF pond	1.99

 Table 21-15:  Breakdown of LOM Mining Operating Costs

Labour	32.4%
Parts and Consumables	67.6%
Diesel Fuel	26.0%
Electricity	0.7%
Tires	16.5%
Repair Parts	13.6%
Explosives	8.4%
Other Parts & Consumables	2.4%

21.5.3 Process Operating Costs

The processing operating cost estimate is divided into the following categories:

· Labour

· Electricity

· Grinding media

· Reagents

· Water supply

· Maintenance materials.

Processing costs for the PEA WOL flowsheet are summarized in Table 21-16.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

 Table 21-16:  Processing Costs for PEA WOL Flowsheet

Area	WOL ($/t milled)
Labour	0.58
Electricity	1.31
Grinding Media and Liners	1.86
Reagents	3.31
Water Supply	0.05
Maintenance Materials	0.49
Total	7.59

Process labour costs were estimated by assuming a work force of 105 personnel, comprising operations, day-crew including tailings management, reagent crew, maintenance, technical, and laboratory staff.  Based on AMEC in-house data, this work force level is consistent with typical 60,000 t/d gold processing plants.  Labour annual salaries were estimated based on AMEC in-house data on similar projects in Western Canada.  The process labour costs were estimated at $12.7 million per year, equivalent to $0.58/t milled.  The personnel and labour cost schedule for processing is shown in Table 21-17.

 Table 21-17:  Processing Labour Costs

Role	Number	Base Salary ($)	Salary Cost – Base $/a	Salary Cost – Loaded $/a
Operators (8/shift)	32	80,000	2,560,000	3,379,200
Operations Supervisors	4	100,000	400,000	528,000
Day Crew/Refinery/Tailings	14	70,000	980,000	1,293,600
Reagent Crew	4	80,000	320,000	422,400
Superintendent	1	180,000	180,000	237,600
Mill Manager	1	200,000	200,000	264,000
Metallurgists	3	120,000	360,000	475,200
Administrator	1	70,000	70,000	92,400
Maintenance Crew	20	90,000	1,800,000	2,376,000
Maintenance Planners/Supervisor	3	100,000	300,000	396,000
Electrical / Instrument Tech	4	90,000	360,000	475,200
Process Control/Electrical Engineer	1	150,000	150,000	198,000
Maintenance Superintendent	1	180,000	180,000	237,600
Laboratory Tech	15	70,000	1,050,000	1,386,000
Chief Chemist	1	100,000	100,000	132,000
Total	105		$9,650,000	12,738,000

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Process power costs were estimated based on estimated operating power loads from the processing equipment for a 60,000 t/d WOL processing plant.  The resulting power costs are summarized in Table 21-18.

 Table 21-18:  Processing Power Costs

Power Costs	MWh/a	$/MWh	Total Power Cost $/a	$/t milled
WOL Operation	644,341	44.40	28,608,745	1.31

Grinding media costs were estimated at $1.86/t milled for WOL operation (Table 21-19).  Reagent costs were estimated at $3.31/t milled for WOL operation (Table 21-20).

 Table 21-19:  Grinding Media Operating Costs – WOL

	g/t use	$/t steel	$/t processed	$/a
Steel Balls – SAG Mill	700	1,200	0.84	18,396,000
Steel Balls – Ball Mill	500	1,400	0.70	15,330,000
Crusher and Mill Liners	80	4,000	0.32	7,008,000
Total				40,734,000

 Table 21-20:  Processing Reagent Use for WOL Circuit

Reagent	g/t use	$/t consumable	$/t processed		$/a
Cyanide	1,000	2,500		2.50	54,750,000
Lime	1,400	150		0.21	4,599,000
CuSO4	32	4,000		0.13	2,833,917
Carbon	80	1,500		0.12	2,628,000
Flocculant	20	6,000		0.12	2,628,000
Diesel (for C reactivation)	-	-		0.10	2,206,519
Other reagents (antiscalant, HCl, NaOH, etc.)	-	-		0.05	1,000,000
SO2	409	200		0.08	1,719,818
Total Reagents					72,437,254

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Water supply costs were calculated by Knight Piésold and include electricity and maintenance for pumping.  Power costs of $610,000/a and pipe and pumping maintenance costs at $380,000/a were estimated.  Total water supply costs are therefore $0.99 million per year.

Maintenance materials are estimated at 2% of direct process plant capital cost, based on in-house AMEC data.  Based on the direct process plant capital cost estimate of $539 million, this results in an annual maintenance materials cost of $10.8 million.

Waste rock disposal and TSF operating costs have been incorporated into mining and processing cost areas.

21.5.4 General and Administrative Operating Costs

The G & A operating cost estimate includes the following:

· G & A labour, including site management, HR and recruitment; safety, environment, accounting, purchasing, warehouse, and logistics and IT.

· Camp and catering contract;

· Insurance;

· Power for camp and administration facilities;

· Freight and logistics costs, other than those already included with process consumables unit costs;

· Site maintenance, including light vehicles and road maintenance; and

· Miscellaneous administration costs, including business travel, office supplies, legal fees, and community relations costs.

G & A costs are estimated at $0.95/t milled, or $20.8 million per year, and are summarized in Table 21-21.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

 Table 21-21:  G & A Operating Costs

Area	Unit Cost ($/t milled)
G & A Labour	0.13
Camp and Catering	0.19
Insurance	0.14
Camp and Admin Power	0.17
Freight and Logistics	0.06
Road and Site Maintenance and Mobile Fleet Maintenance	0.14
Power Line Maintenance	0.03
Other G & A	0.10
Total	0.95

21.5.5 Operating Cost Summary

The average LOM operating cost for the Project is estimated at $15.17/t milled, and $13.01/t milled after silver credit (Table 21-22).  After accounting for royalty and a silver credit of US$90/oz Au produced, the average LOM operating cost is US$543/oz Au produced.  The cash costs/oz Au produced during periods of mine operation are shown in Table 21-23.

 Table 21-22:  Summary of Blackwater LOM Operating Costs by Cost Area

	LOM Unit Cost
Area	($/t milled)	($/oz Au produced)	(US$/oz Au produced)
Mining	6.21	276	259
Processing	7.59	337	317
G&A	0.95	42	40
Refining Costs	0.23	10	9
Royalties (0.56%)	0.18	8	8
Total Cash Costs	15.17	674	633
Silver Credits	(2.16)	(96)	(90)
Total Cash Costs incl. silver credit and royalties	13.01	578	543

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

 Table 21-23:  Cash Costs per Ounce of Gold Produced During Periods of Mine Operation

	Years			LOM
	1–5	1–15	16–17
Cash Costs - $/oz	579	666	811	674
Cash Costs - $/oz inc. Ag credit, royalty	497	570	722	578
Cash Costs - US$/oz	544	626	762	633
Cash Costs - US$/oz inc. Ag credit, royalty	467	536	678	543

21.6 Comments on Section 21

In the opinion of the QPs, the following conclusions and comments are made:

· Capital and operating costs were prepared according to each individual consultant’s area of expertise

· Total capital costs are estimated at $1,814 million.  The estimate was based on a combination of semi-detail estimating for some components of the Project and capacity factor or equipment factor estimating for others.  The estimate excludes off-site facilities, such as an off-site reagent transfer and storage facility.

· The LOM average operating cost is estimated at $15.17/t milled.  Operating costs include mining, processing, G & A, royalties, refining charges, and product transport costs.

· After accounting for a silver credit of US$90/oz Au produced, the average LOM operating cost is US$543/oz Au produced.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

22.0 ECONOMIC ANALYSIS

The following section is preliminary in nature and partly based on Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the PEA based on these Mineral Resources will be realized.  Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.  Dates cited are for illustrative purposes only, as a decision to proceed with mine construction will require formal approvals from appropriate regulatory authorities and from New Gold’s Board, and and will require support from additional, more detailed, studies, and declaration of Mineral Reserves.

The results of the economic analyses discussed in this section represent forward-looking information as defined under Canadian securities law.  The results depend on inputs that are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here.  Information that is forward-looking includes:

· Mineral Resource estimates;

· Assumed commodity prices and exchange rates;

· The proposed mine production plan;

· Projected recovery rates;

· Infrastructure construction costs and proposed operating costs;

· Project development schedules;

· Assumptions that New Gold will acquire the social licence to operate; and

· Assumptions that an EA will be approved by Provincial and Federal authorities.

22.1 Methodology Used

A discounted cash flow (DCF) model was developed to assess the value of the Project throughout the PEA phase, where different project configurations were evaluated with the objective of selecting a base case.

The DCF model developed assumes 100% equity financing (no debt), and all figures are expressed in second quarter (Q2) 2012 constant dollars or real terms.  No inflation or real cost escalation was assumed.  Cash flows are assumed to occur mid-period.  All monetary figures are expressed in Canadian dollars unless otherwise specified.

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The Project value is determined on a pre-tax basis and discounting was performed to the start of 2015, which marks the first year of proposed project construction.  Therefore, expenditures incurred prior to 2015 are considered sunk and are not considered in the analysis.  These expenditures cover additional exploration and project studies.

A post-tax DCF model was also developed using the taxation assumptions stated in Section 22.2.2.

Mining pre-production is planned to start in 2016 and 2017 would mark the first year of revenue.

22.2 Financial Model Parameters

22.2.1 Assumptions

The DCF model is based on the underlying economic assumptions listed in Table 22-1.  The economic model includes royalties payable on Claims 515809 and 515810, as shown in Table 22-2.  It also assumes that the option to buy out the royalty on claims 509273, 509274, and 509275 for $2 M will be exercised in 2016, as this buyout price is substantially less than the NSR royalty on these claims.

The construction period of the Project is estimated to be two years and the major capital outlays are therefore scheduled to occur during the two years prior to production, 2015 and 2016, with the first year of production being 2017.  The initial capital cost is $1,814 million.

The mine production schedule and all mining-related costs (mining material to be processed, stockpiled, or treated as waste) were provided on a per unit basis by IMC.  Over the LOM, mining costs are $1.90/t mined.  Processing costs are estimated at $7.59 per tonne milled.  General and administrative costs are variable and are estimated at $0.95 per tonne milled.  Total operating costs, including refining costs, over the LOM are summarized in Table 22-3.

Sustaining capital will be split among three areas:  TSF, mining, and plant.  The LOM costs and annual averages for each area are listed in Table 22-4.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

 Table 22-1:  Economic Inputs

Metal Price
Gold	US$/oz	1,275
Silver	US$/oz	22.5
Discount rate	%	5.0
Exchange rate (1 CAD = X USD)	CAD/USD	0.94
Refinery Deductions
Payable factor – Au	%	99.9
Payable factor – Ag	%	95.0
Dore refining treatment charge	$/oz Au + Ag	0.80
Au transport and insurance	$/oz Au	0.32

 Table 22-2:  Royalty Requirements Covering Blackwater Deposit

Claim Number	Claim Name	Remaining NSR Royalty	Remaining Buy out Amount
509273, 509274, 509275	Davidson	2%	$2,000,000
515809	Dave	1.5%	-
515810	Jarrit	1%	-

 Table 22-3:  Operating Costs Statistics (Unit Costs Averaged over the Mine Life)

Operating Costs	Units	LOM Total	LOM $ per Tonne Milled	LOM US$ per ounce Produced
Mining Cost	M$	2,212	6.21	259
Processing Cost	M$	2,703	7.59	317
Management and Admin	M$	338	0.95	40
Refining Costs	M$	81	0.23	9
Royalty	M$	65	0.18	8
Total	M$	5,398	15.17	633
Silver Credit	M$	(769)	(2.16)	(90)
Costs Net of Silver Credit	M$	4,637	13.01	543

 Table 22-4:  Sustaining Capital Costs per Mine Operating Area

Mine Operating Area	LOM Total M$	LOM Annual Average M$/year
Mining	288.8	17.0
Plant	65.0	3.8
Tailings Storage Facility	217.7	12.8
Total Mining Operating Expense	571.5	33.6

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

22.2.2 Taxes

The tax rate calculations and taxation assumptions used in the after-tax financial model are summarized below.

The federal government imposes income tax on mining income at the same rate that applies to other types of income. Commencing January 1, 2012 the federal rate applicable to resource profits is 15%. British Columbia’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 income tax liability in respect of the Blackwater Project.

Under the British Columbia Mineral Tax Act mining taxes are imposed mine by mine in two stages, namely

· a 2% tax on net current proceeds

· a 13% tax on net revenue.

The initial 2% tax is a form of minimum tax, which is deductible in full against the 13% tax.  Both the net current proceeds tax and the net revenue tax have been factored in the Blackwater life of mine model

The federal and provincial tax legislation provide 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. As these deductions and allowances are only available when incurred, a high-level assumption was made with regards to the allocation of expenditures between the three categories in the life-of-mine model.

22.2.3 Working Capital

Detailed analysis of working capital requirements was not performed for the purpose of this study; however, working capital allowances required to operate this Project are not expected to have a significant impact on the cash flow of the mine.

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22.3 Pre-Tax Financial Results

The calculated pre-tax NPV5 for the base case (NPV5%, and 2015) is $1,872 million.  After conversion to US$, the pre-tax NPV5 for the base case is US$1,655 million, the IRR is 16.4% and the simple payback period is 4.7 years from the start of production.

Average life of mine cash costs, net of royalties are US$633/oz, and net of royalties and silver credits, US$543/oz.

Figure 22-1 shows gold annual gold production and cash costs (net of silver credits).  A production summary table is shown in Table 22-5.  The production costs and cash flows on an annualized basis are included in Table 22-6.  Project value creation is shown in Figure 22-2 depicting the annual net cash flow by year, and a cumulative discounted total.

 Figure 22-1:  Gold Production and US$/oz Cash Costs (net of royalty and silver credits)

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

 Table 22-5:  Production Summary

	Units	LOM Total	LOM Annual Average	Year 1–5 Annual Average	Year 1–15 Annual Average	Year 16–17 Annual Average
Life of Mine1	Years	16.4	NA
Metal Price
Gold	US$/oz	NA	1,275
Silver	US$/oz	NA	22.5
Production
Mineralized Material Milled - Indicated	kt	255,935	15,606	17,239	15,764	13,913
Mineralized Material Milled - Inferred	kt	99,992	6,097	4,223	5,990	7,242
Waste Mined	kt	839,102	51,165	61,040	59,073	14,157
Strip Ratio W:MM	NA	2.36		2.42	2.72	0.66
Gold grade - Indicated	g/t	0.87		1.00	0.90	0.49
Gold grade - Inferred	g/t	4.3		5.4	4.4	3.3
Silver grade - Indicated	g/t	0.65		0.75	0.66	0.54
Silver grade - Inferred	g/t	7.9		4.7	8.5	3.1
Metal Feed
Gold - Indicated	koz	7,172
Gold - Inferred	koz	2,082
Silver - Indicated	koz	35,276
Silver - Inferred	koz	25,508
Process Recoveries
Gold	%	86.6
Silver	%	52.9
Metal Recovered
Gold - Indicated	koz	6,222
Gold - Inferred	koz	1,791
Silver - Indicated	koz	18,648
Silver - Inferred	koz	13,478
Gold - Annual production rate	koz		489	569	507	296
Silver - Annual production rate	koz		1,959	1,966	2,039	1,103
LOM Recovered Metal Value
Gold	M$	10,868
Silver	M$	769

1 Excludes pre-production year

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

 Figure 22-2:  Annual Cash Flows and Cumulative Project Value (CDN$)

22.4 Post-Tax Financial Results

Allowing for taxes, the post-tax NPV5 for the base case is US$1,123 million, the IRR is 14.0% and the simple payback period is 4.8 years.

22.5 Sensitivity Analysis

Sensitivity analysis was performed on the Project to gauge its robustness against favourable and unfavourable changes to project variables.

Figure 22-3 demonstrates how the project pre-tax NPV is influenced by the following variables:  gold price, exchange rate, operating expenses, and capital expenses.  The steepness of the lines indicate the Project is more exposed to swings in the gold price and the USD:CAD exchange rate, than to changes in capital or operating costs.  The change in the gold price mirrors changes in deposit gold grades.

Table 22-7 includes the sensitivity figures depicted on the chart above plus sensitivities to mining and processing costs, and silver price pre-tax.  Table 22-8 shows the sensitivity of economic performance to alternative higher gold and silver prices, and respective exchange rate assumptions.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

 Figure 22-3:  Selected Sensitivities: Major Dirvers Pre-Tax NPV (CDN$)

 Table 22-7:  Sensitivities, Pre-Tax NPVat 5% Discount Rate (CDN$)

	-30%	-20%	-10%	0%	10%	20%	30%
Capital expenditure	2,519	2,304	2,088	1,872	1,656	1,440	1,225
Operating expenditure	2,863	2,533	2,202	1,872	1,542	1,211	881
Gold price	(198)	492	1,182	1,872	2,562	3,252	3,942
Silver price	1,736	1,781	1,827	1,872	1,917	1,963	2,008
Foreign Exchange Rate	5,024	3,711	2,689	1,872	1,203	646	175
Mining cost	2,295	2,154	2,013	1,872	1,731	1,590	1,449
Processing Cost	2,377	2,208	2,040	1,872	1,704	1,536	1,367

 Table 22-8:  Sensitivities of Economic Performance vs. Metal Prices and Exchange Rates

Summary of Project Economics
Gold Price (US$ per ounce)	Silver Price (US$ per ounce)	US$/CDN$ foreign exchange	5% NPV (US$ billions)			IRR (%)			Payback Period (Years)
			Pre-Tax	After-tax		Pre-tax	After-tax		Pre-tax	After-tax
1,275	22.50	0.94	1,655	1,123		16.4	14.0		4.7	4.8
1,600	30.00	0.97	3,338	2,246		25.9	22.0		3.0	3.1
1,775	34.50	1.00	4,202	2,815		30.4	25.8		2.6	2.7
1,800	35.00	1.00	4,338	2,905		31.1	26.4		2.5	2.6

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22.6 Comments on Section 22

In the opinion of the QPs, the Project as outlined in this Peport, based on Indicated and Inferred Mineral Resources, and using the assumptions in the Report, displays positive project economics.  The calculated pre-tax NPV5 for the base case is $1,872 million.  After conversion to US$, the pre-tax NPV5 for the base case is US$1,655 million, the IRR is 16.4% and the simple payback (undiscounted) period is 4.7 years from the start of production.

The Project is most sensitive to changes in the gold price, then changes in the USD to CAD exchange rate, and less sensitive to changes in the operating cost.  It is least sensitive to capital cost changes.

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23.0 ADJACENT PROPERTIES

This section is not relevant to this Report.

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24.0 OTHER RELEVANT DATA AND INFORMATION

New Gold has commenced an assessment of risks and opportunities for the Project based on likelihood and consequence of outcomes.  Many of the risks are associated with current uncertainties related to the limited testing and technical information about the Mineral Resource estimate, material properties, and metallurgical parameters, as would be expected in a PEA.  Plans have been recommended in relevant geological, geotechnical, mining, metallurgy, and environmental areas as appropriate to better quantify aspects of the Project during more detailed project studies.

24.1 Risks

Some of the key risks that have currently been identified and are being investigated are outlined below.

24.1.1 Conversion of Mineral Resources to Mineral Reserves

There is a risk that some or all of the Inferred Mineral Resources may not be able to be converted to higher-confidence Mineral Resource categories, or eventually to Mineral Reserves.  There is also a risk that some of the Indicated Mineral Resources may not support full conversion to Mineral Reserves.  This risk is planned to be mitigated through completion of infill drilling and additional metallurgical testwork.

24.1.2 Processing

Some uncertainty remains on the metal recoveries.  It is possible that gold and silver recoveries may be lower on some material types (low-grade stockpile and high-silver GS10 zone).  These risks will be mitigated through more metallurgical testwork and additional process design.

24.1.3 Waste Management

The waste management plan takes into account the collection and containment of contact water from waste materials.  There is some uncertainty about the amounts and exact leaching characteristics of waste materials; this is being studied further through the ARD assessments.  If necessary, changes will be made in the segregation and placement of materials—in embankments, for example—and in the design of waste facilities to manage these risks.

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24.1.4 Project Economics

A variety of economic risks are associated with the Project, including:

· Change in metal prices and exchange rate

· Change in input costs, primarily labour, fuel, and bulk materials for mining and processing

· Escalation in capital cost

· Project schedule delay

Metal prices, exchange rates, and input costs are generally beyond New Gold’s control.  Escalation in capital costs or possible delays in the construction schedule could result from shortages of a skilled and productive construction workforce.  Schedule delays could also result from delays in receipt of construction permits.  Skilled labour shortages could furthermore translate into the operating phase of the Project, increasing operating costs.  Contracting, recruitment and retention strategies will be developed to minimize these risks.

The effect of some of these risks on the Project economic assumptions can be assessed in the sensitivity analysis.

24.2 Opportunities

A number of key opportunities have been identified that could increase project value.

24.2.1 Delineation of Additional Mineralization

Drilling completed since the database close-out date for the Mineral Resource estimate has shown the potential for additional mineralization to the north and at depth.  New Gold continues to actively drill the deposit to define the deposit limits.  Discovery of additional mineralization that can support Mineral Resource estimates and conversion of all or part of these Mineral Resource estimates to Mineral Reserves has the potential to add project value by extending mine life and increasing throughput.

24.2.2 Pit Slopes

The current pit wall slopes are based on limited geotechnical data.  If the pit slopes can be steepened, then additional mineralization may come within the boundary of the LG pit shell and may be able to be classified as Mineral Resources.  In addition, the stripping ratio would decrease because of the lower volumes of waste that would need to be removed on the margins of the pit.  Further geotechnical drilling and analysis are planned.

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24.2.3 Mining

Mining costs may be reduced by mechanical waste conveyance and by optimizing mine plan, stockpile locations, and mining equipment selection.

24.2.4 Process Flowsheet and Parameters

Reducing process costs and increasing gold and silver recoveries may be achieved by optimizing the process flowsheet, grind sizes, cyanide consumption, and leach tank capacity.  These opportunities will be investigated.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

25.0 INTERPRETATION AND CONCLUSIONS

The QPs have reached the following conclusions and made the following interpretations as a result of the review of the PEA document:

· Information provided by legal experts and New Gold support New Gold’s ownership claims to the Project mineral tenure.

· The Project consists of 69 mineral claims.  A total of 66 of the claims are recorded in the name of New Gold.  Three claims, 637203, 637205, and 637206, termed the JR Claims, are recorded in the name of third parties.  New Gold was granted an option to acquire a 100% interest in the JR Claims, subject to a royalty.  No opinion was provided as to unrecorded interests in or affecting the claims as these cannot be ascertained through inquiries with the Gold Commissioner’s office.

· The currently-defined Blackwater deposit spans the one of the three Davidson claims (509273 but not 509274 and 509275), the Dave claim (515809) and the Jarrit claim (515810).

· Future development will require conversion of the mineral claims covering the Blackwater deposit to mining leases.

· The Claims are located on Crown land.  Limited searches that have been conducted did not reveal any dispositions of surface rights by the Crown under the Land Act within the Project area.  A number of parties hold usage rights, including trap lines, forestry cutting permits and a communications reserve/notation..  There is a reasonable expectation that with appropriate supporting studies as required by permitting authorities that land access and provision of land for infrastructure development for any proposed mining activity will be achievable.

· Mining activities should be capable of being conducted year-round.

· There is sufficient suitable land available for any future tailings disposal, mine waste disposal, and related mine infrastructure within the mineral claims.

· Five royalties are payable:  a 2% NSR on mineral claims 509273, 509274, and 509275; a 1.5% NSR on mineral claim 515809, a 1% NSR on mineral claim 515810, a 3% NSR on mineral claims 637203, 637205, and 637206, and a 2% NSR on mineral claim 835014.  New Gold may purchase the entire 2% NSR on mineral claims 509273, 509274, and 509275 for $2 M, 2% NSR of the 3% NSR on mineral claims 637203, 637205, and 637206 for $1 M (at any time after it exercises the option to acquire a 100% interest in those claims), and 1% NSR of the 2% NSR on mineral claim 835014 for $1 M.

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· The economic analysis assumes that the royalty on mineral claims 509273, 509274, and 509275 is bought out, and that the only two remaining royalties that affect the Blackwater deposit will be a 1.5% NSR on mineral claim 515809, and a 1% NSR on mineral claim 515810.

· Knowledge of the deposit settings, lithologies, and structural and alteration controls on mineralization, and the mineralization style and setting is sufficient to support Mineral Resource estimation.  The deposit is considered to be an example of a low to intermediate epithermal system.

· The exploration programs completed to date are appropriate to the style of the deposits and prospects within the Project.  The exploration and research work supports the genetic and affinity interpretations.

· The quantity and quality of the lithological, geotechnical, and collar and down hole survey data collected in the exploration and delineation drill programs completed from 2009 to 2012 are sufficient to support Mineral Resource estimation.  Data from holes drilled between 1981 and 1994 have no documented QA/QC information, and they are not deemed acceptable for use in resource estimation.

· Sampling methods are acceptable, meet industry-standard practice, and are acceptable for Mineral Resource estimation.

· The quality of the gold and silver analytical data are sufficiently reliable to support Mineral Resource estimation and sample preparation, analysis, and security are generally performed in accordance with exploration best practices and industry standards.

· Gold analyses are performed by fire assay/AA finish methods and silver analyses are performed by ICP.  Silver ICP analyses are not known with the same precision and do not have the same quality control support as gold fire assay analyses.

· New Gold is currently investigating a possible decrease in analytical precision noted in analytical QA/QC samples analyzed by ALS since July 2011.

· The process of data verification performed by the QPs indicates that the data collected from the Project during the 2009 to 2012 work programs adequately reflect deposit dimensions, true widths of mineralization, and the style of the deposits, and adequately support the geological interpretations, and the analytical and database quality.

· QA/QC with respect to the results received to date for the 2009, 2010, and 2011 exploration programs is acceptable, and protocols have been well documented.  A lower level of QA/QC exists for the seven drill holes completed between 2005 and 2006, but the data are considered to be sufficiently reliable to support a resource estimate.

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· Drill data are typically verified prior to Mineral Resource estimation by comparing data in the Project database to data in original sources.  For most of the data, the original sources are electronic data files; therefore, the majority of the comparisons were performed using software tools.  No significant errors were found with the database that would preclude use in Mineral Resource estimation.

· Metallurgical testwork and associated analytical procedures were performed by recognized testing facilities, and the tests performed were appropriate to the mineralization type

· Overall results indicate that the gold in Blackwater mineralization is free-milling (non-refractory).  A portion of the gold, between 17% and 32%, liberates during grinding and is amenable to gravity concentration.  In addition, a preliminary diagnostic leach test indicated that 90% to 95% of the overall gold is cyanide-soluble, with only 1% to 5% of the gold occluded in sulphides.  Silver, despite higher grades, exhibited poor amenability to low mass pull gravity concentration, and diagnostic leach tests showed moderate dissolution, in the range of 55% to 62%.  This level of dissolution is in line with observations of silver minerals such as friebergite, which has been reported to be refractory

· Estimations of Mineral Resources for the Project conform to industry best practices, and meet the requirements of CIM (2010).  A total of 417 holes totalling 147,619 m were used to support the grade estimations.  Due to lack of QA/QC and accurate survey information, holes drilled before 2009 were not used for statistical analysis or grade estimation of the PEA Mineral Resource, but were used in forming the lithological wire frame construction.  The Mineral Resource estimate is based on reasonable assumptions of economic extraction assuming open pit mining methods.  Assumptions include US$1,300/oz Au, US$24/oz Ag, 83.6% Au recovery, 44.9% Ag recovery, $1.52/tonne mining cost, $1.90/tonne waste mining cost, $10.52/tonne process and G&A cost, and average pit slope angles of 40°.

· Indicated Mineral Resources at a gold equivalent cut-off grade of 0.3 g/t AuEq total 267 Mt grading 0.876 g/t Au and 4.3 g/t Ag.  In addition, there is an Inferred Mineral Resource of 120 Mt grading 0.687 g/t Au and 7.3 g/t Ag at a cut-off grade of 0.3 g/t AuEq.

· The distribution of resource tonnes by material type is 3% within the oxide zone, 9% within the transition zone, and 88% in the sulphide zone.

· To support declaration of higher confidence Mineral Resource categories, the following will need to be addressed:

- additional quality control support for silver results;

- silver metallurgical recovery results for high Ag-low Au zone modelled as domain GS-10; and

- additional infill drilling.

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· Factors which could affect the Mineral Resources are:

- commodity price assumptions;

- assumptions that all required permits will be forthcoming;

- pit slope angles;

- metal recovery assumptions; and

- mining and process cost assumptions.

- relogging of older drill holes may result in local changes to the lithological interpretations, on which the current block model, and resource estimate, is based.

· Drilling is ongoing.  A review of the drill data available as at 28 August 2012 was performed to compare the results with the block model.  Although incorporation of these drill holes in an updated block model will result in local differences in interpretation of the geological and grade shells, visual inspection of the drill holes against the block model indicated the results were generally consistent with the current interpretations as used in the model.

· The open pit mining method will use conventional mining techniques and be based on a truck-and-shovel operation.

· The proposed open pit will be mined in four phases.  Pit design has appropriately considered available geotechnical information.  Mine design has included consideration of the high rainfall and snowfall conditions.  Geohazard assessments will be required.

· The mine plan is preliminary in nature and partly based on Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the PEA based on these Mineral Resources will be realized.  Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.  The mine plan includes both Indicated and Inferred Resources to make up the mill feed tonnage.  Material not processed is considered to be waste.

· The LOM production will be 255.9 million tonnes (Mt) from the Indicated category and 100.0 Mt from the Inferred category, and mill feed will be delivered to a primary crusher that feeds a WOL gold–silver recovery plant.  The milling rate is 60,000 t/d, or 21.9 Mt/a.  The mine life is 16.4 years after commercial production begins.  The total tonnage mined from the open pit is 1,195 Mt at an initial daily mining rate of 169,000 t/d and increasing to 273,000 t/d during Years 3 through 10 before declining after the peak waste mining period.  A total of 7.2 million ounces of gold and 35.3 million ounces of silver are estimated to be contained in the Indicated category of mill feed tonnage during the life of the mine.  An additional 2.1 million ounces of gold and 25.5 million ounces of silver are estimated to be contained in the Inferred category of mill feed.

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· Some of the mill feed, referred to as low-grade mineralized material, will be stockpiled in the early years for re-handle into the mill at the end of the mine life.  Another sub-set of the mill feed is within a high silver–low gold area in the northern part of the deposit, referred to as high silver mill feed; this material will also be stockpiled and re-handled as mill feed.

· Five types of waste rock have been identified and the mine design appropriately incorporates consideration of these materials.  The waste storage facilities are designed to accommodate the waste rock production over the LOM.

· Equipment and personnel requirements have been considered in relation to the mine plan and are appropriate for the mine production rate, sequencing, and mine life.

· The Project will use conventional mineral processing equipment to produce doré.  The plant design is based on conventional technology and proven equipment.

· The process design is based on the metallurgical testwork and is appropriate to the grind, flotation and recovery characteristics defined for the different mineralization types.

· Run-of-mine mineralized material from the open pit will be crushed and conveyed to the concentrator where the material will be ground to allow gold extraction and recovery.  Some gravity-recoverable gold will be recovered from processing a bleed stream within the grinding circuit.

· During initial operations, the plant will be processing primarily oxide and transition materials, which are softer, and therefore throughputs as high as 70,000 t/d of transition mineralization and 85,000 t/d of oxide material could potentially be achieved.

· Mill feed ground to a P80 of 150 µm will be thickened ahead of leaching to optimize slurry density and leached in a conventional cyanide leach circuit.  Gold will be recovered from loaded carbon, electrowon, and smelted into doré bullion.  CIP residue will be thickened and treated in the cyanide destruction circuit prior to disposal in the TSF.

· Trade-off studies within the PEA WOL flowsheet as described in this Report should be investigated to minimize capital and operating costs.  Trade-off studies should be investigated to minimize capital cost while addressing the potentially challenging flotation performance on oxide and transition mineralized material.

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· The WOL process configuration was selected for the current PEA based on the available metallurgical data as of May 2012.  The WOL, FCL, and Hybrid process options need further investigation to confirm which is optimal from an economic, operational, and environmental perspective.

· The Project is a greenfields site and will require construction of infrastructure to support the planned producing facilities.

· Tailings and waste facility design is based on appropriate geotechnical testwork, and acid-base accounting tests.

· A number of water control structures are planned, and will include the collection and diversion of surface water runoff, sediment control ponds, and pump-back systems.

· Process and potable water will be supplied via a 15.7 km long pipeline from Tatelkuz Lake.

· Road access is expected to be via the existing Kluskus Forest Service Road.

· Employees are expected to live in several local and regional communities once operations have commenced.

· New Gold is expected to be able to market gold produced from the Project. Sales contracts that could be negotiated would be expected to be within industry norms. However, the majority of production would be expected to be spot marketed.

· Baseline environmental studies have been initiated and are ongoing.

· Dissolved aluminum and total cadmium arising from natural sources exceeded guidelines at a number of sites.

· Culturally modified trees were identified in one location near the Project area.  No other archaeological sites, historic heritage sites, or evidence of any previous archaeological investigations were identified within the project area

· The Project is subject to review under the Environmental Assessment Act (British Columbia) and the Canadian Environmental Assessment Act (Canada).  A Project Description to initiate the EA process, has been written in parallel with completion of the PEA, and is due for submission in late 2012.  Following EA approval, additional permits, licences, authorizations, and certificates will be required to proceed to construction of the Project.  New Gold has identified and commenced planning for obtaining all such approvals in parallel with the EA approval process.

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· The Project is subject to federal and provincial permitting requirements, which include public and First Nations consultations.  The BCEAA and CEAA, 2012 have specific, regulated timelines, and in addition, provincial and federal permits have specific timelines and requirements that must be met.  New Gold has outlined a permitting schedule based on these timelines; however changes to the schedule may occur as a result of the consultation processes and any subsequent amendments that may be then required to support furtherance of permit application processes.

· Existing environmental liabilities are related to the exploration activities that New Gold has undertaken to date or to activities conducted by third parties prior to New Gold’s acquisition of the Blackwater mineral claims.  The current established security deposit for project reclamation is approximately $1 million; New Gold is of the opinion that currently-existing liabilities are within the reclamation bond amount.

· New Gold has made contact with First Nations groups who have interests in the Project area.

· New Gold is engaged in ongoing consultation with provincial and federal ministries and agencies about research design, permits, and environmental assessment questions and issues as they arise.

· Capital and operating costs were prepared according to each individual consultant’s area of expertise.

· Total capital costs are estimated at $1,814 million.  The estimate was based on a combination of semi-detail estimating for some components of the Project and capacity factor or equipment factor estimating for others.  The estimate excludes off-site facilities, such as an off-site reagent transfer and storage facility.

· Mine closure costs are estimated at $101 million.

· It is assumed that $77 million salvage value can be obtained from major mine equipment and the crushing and grinding equipment from the processing plant.

· A diesel cost of $1.04/L has been assumed, based on an average oil price of US$90/bbl and exchange rate of CAD $1 = US$ 0.94.

· An average electricity tariff rate of 4.44 c/kWh was assumed for Blackwater, based on the BC Hydro tariff rate effective at May 2011, Schedule 1823: Stepped Rate, BC Zone 1 (BC Hydro, 2012).

· The LOM average operating cost is estimated at $15.17/t milled.  Operating costs include mining, processing, and general and administration (G & A) costs.  Costs exclude royalties, refining charges, and product transport costs.  After accounting for a silver credit of US$90/oz Au produced, and royalty, the average LOM operating cost is US$543/oz Au produced.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

· For mining, the hourly personnel head count begins during preproduction with 89 in operations and 59 in maintenance.  The number of hourly persons peaks in Year 9 with 237 in operations and 146 in maintenance.  In addition, 40 salary staff will be required in the mine and 33 salary staff will be required for G & A duties.  Process personnel assumptions comprise a work force of 105 personnel, consisting of operations, day-crew including tailings management, reagent crew, maintenance, technical, and laboratory staff.

· The economic analysis is preliminary in nature and partly based on Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the PEA based on these Mineral Resources will be realized.  Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

· Discounted cash flow (DCF) models were developed to assess the value of the Project, pre-tax and after-tax.  The construction period of the Project is estimated to be two years and the major capital outlays are therefore scheduled to occur during the two years prior to production, 2015 and 2016, with the first year of production being 2017.

· The calculated pre-tax NPV5 from 2015 for the base case is $1,872 million.  After conversion to US$, the pre-tax NPV5 for the base case is US$1,655 million, the IRR is 16.4 % and the simple payback (undiscounted) period is 4.7 years from the start of production.  Allowing for taxes, the post-tax NPV5 for the base case is US$1,123 million, the IRR is 14.0% and the simple payback period is 4.8 years.

· The Project is most sensitive to changes in the gold price, then changes in the USD to CAD exchange rate, and less sensitive to changes in the operating cost.  It is least sensitive to capital cost changes.

· New Gold has performed a risk and opportunity analysis.  Risks identified include conversion of some or all of the Inferred mineral resources to higher-confidence mineral resource categories, process recoveries, amounts and exact leaching characteristics of waste materials, and cost escalations due to various factors.  Opportunities identified include delineation of additional mineralization that can support mineral resource estimates, steepening of pit walls if geotechnical conditions permit, and reductions in mining and processing costs.

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

26.0 RECOMMENDATIONS

AMEC recommends that New Gold) considers the recommendations in this section as activities which will support Project advancement should New Gold determine that a Feasibility Study (FS) is warranted.

As part of the recommended work program, the following areas of work should be considered: additional drilling, topographic surveys, geotechnical studies, engineering and metallurgical studies, land management, including applications for mining leases where appropriate, additional baseline studies, and environmental and permitting activities. Further areas for work are also likely to be identified as activities progress.

The program is designed as a two-phase program, with all elements of the first phase of the program conducted concurrently; this work phase is in progress. Phase 1 work comprises various testwork programs. Phase 2 comprises a number of studies that will culminate in the completion of a FS, and will allow New Gold to submit an EA and EIS.

AMEC’s recommendations do not include provision for construction activities for site and access infrastructure such as the road.

The Phase 1 activities include infill and delineation drilling and assaying, data collection, metallurgical testing, engineering optimization studies and investigations and studies and activities to support the basis of the FS and public consultation processes. Some more specific recommendations for work focus have also been included for mineral resource estimation, road and powerline design, and plant design purposes. The total cost of these activities is estimated to be between about $90 M to $102 M.

The Phase 2 activity comprises completion of a Feasibility Study, an EA, an EIS, and on-going community and government engagement activities estimated at between about  $38 M and $43 M.

Total program costs are likely to range between approximately $128 M and $145 M.

AMEC notes that the Phase 1 work is already in progress, which includes infill, delineation and condemnation drilling and assaying, geotechnical drilling and testing for both the planned TSFs and the open pit, metallurgical sample collection, assaying and testing and discussions relating to road usage and grid power connection.

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26.1 Phase 1 - Testwork

The Phase 1 activities include infill and delineation drilling and assaying, data collection, metallurgical testing. engineering optimization studies and investigations and studies and activities to support the basis of the FS. Some more specific recommendations for work focus have also been included for mineral resource estimation, road and powerline design, and plant design purposes. The total cost of these activities is estimated to be between about $90 M and $102 M. The most significant cost contributors will be the drill programs and logistical support required for this work.

26.1.1 Drill Programs

Exploration activities at Blackwater project include infill and delineation drilling to support potential upgrade of classification categories to Measured & Indicated, condemnation drilling of key site facilities, waste rock characterization and reconnaissance exploration of the greater Blackwater project area. Total drilling costs are estimated at $54 M to $61 M.

· Infill and delineation: the program to support potential upgrade of blocks in the Blackwater mineral resource to Measured or Indicated classification will involve approximately 400 vertical and angled holes totaling approximately 110,000 m. Drilling will be done at a nominal drill hole spacing ranging from 25 to 50 m, and will continue through 2012 and into 2013 as required. The total estimated cost to complete the delineation and infill drilling program is approximately $31 M to $35 M.

· Condemnation drilling:  the condemnation drilling program is designed to test and confirm the mineralisation potential of areas selected for site facilities and mine infrastructure. Ideally these areas will be confirmed to be devoid of mineralisation and therefore suitable for locating future site facilities and infrastructure.  The condemnation drilling program will involve approximately 125 vertical holes totaling 50,000 m at nominal 500 m grid spacing and drilled to an average depth of 400 m. The condemnation drilling program will continue through the end of 2012 and into 2013 as required. The total estimated cost to complete the condemnation program is approximately $13 M to $15 M.

· Waste Characterization Program: the waste characterization program involves approximately 10 diamond drill holes totaling 4,000 m. Drill holes will be directed and inclined toward the ultimate high wall of the proposed open pit. It is designed to obtain representative samples of the various types of waste rock material which will be analysed to determine the ARD characteristics of potential waste material in order to support the design criteria for future waste rock containment facilities.  The waste characterization program will continue through the end of 2012 and into 2013 as required. The total amount required to complete the ARD program is approximately $1 M.

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· Blackwater Area Exploration: this program involves a combination of geologic mapping, geochemical sampling and geophysical surveys to explore and define areas with potential for the discovery of additional mineralisation that could potentially be included in a mine plan with appropriate supporting estimates and studies. The Blackwater area exploration program will continue through 2012 and into 2013 as required. The total estimated cost to complete this program is approximately $5 M.

· Geotechnical drilling to support the design of plant site facilities, open pit and TSF is required.  The work programs have been designed to enhance the understanding of the geotechnical conditions onsite and include the following: 36 sonic drillholes,    14.2 km of seismic refraction surveys, surficial geology assessment of the Davidson drainage and a geotechnical laboratory program to characterize materials for use in construction. Additional drilling and testing will be undertaken in the open pit area to determine hydrogeological conditions, approximately 1,600 m of 250 mm diameter and 4,800 m of PQ drilling, and including consultant support for programs (estimated cost of $3 M to $4 M).

· Drilling for additional metallurgical testing, to complement the samples already selected for the FS metallurgical program. The program is estimated to comprise approximately 2,000 m of drilling (estimated cost of $1 M).

26.1.2 Camp Operations

· Due to the location of the Project, significant provision is required for site logistics including camp support, site management, provision of core laydown areas, site equipment maintenance, communications and helicopter and helicopter fuel costs to support data collection activities and any Feasibility Study activities. In addition, provision for site personnel and overheads is included (estimated cost of $15 M to $17 M).

26.1.3 Administration Costs

· The administration costs for the project include the labour costs of the project management team, tax, legal, IT, offices etc (estimated cost of $8 M to $9 M).

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New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

26.1.4 PEA Engineering Studies

· The costs to complete the 43-101 Technical Report following the PEA are estimated at $1 M.

26.1.5 Metallurgical Testwork

· An extensive program of metallurgical testing is in progress to investigate recovery parameters for WOL and FCL flowsheets, grind size, cyanide consumption, leach and flotation retention times, tailings cyanide destruction, gravity recovery and the variability of metal recoveries over the deposit. Testwork is currently underway at McLelland laboratory and Dawson laboratory, and will be also be conducted at SGS Lakefield. Additional work is planned for further grinding and recovery characterization (estimated cost of $1 M to $2 M).

26.1.6 Engineering Optimization Studies

· A number of trade-off studies and design updates will be required, including mining strategy optimization, mine bench height and equipment studies, PAG waste material handling study, mine phasing and production planning, grinding circuit configuration study, cyanide destruction reagent study, CIP vs pumpcell study, updates and review of tailings design and water and waste management plans, and process flowsheet selection study (estimated cost of $10 M to $12 M).

26.2 Phase 2 - Studies

The information collected as part of the Phase 1 work program will be used to complete a Feasibility Study that will incorporate an updated Mineral Resource and Mineral Reserve estimate, mine plan, infrastructure layout and economic analysis, and to complete an Environmental Impact Statement. The work is estimated to cost at a minimum of between $38 M and $43 M.

26.2.1 Feasibility Studies

· Feasibility study-level engineering and costing for the mine, plant and all associated infrastructure will be required to complete the FS, with a number of specialist consultants providing design of specific areas (estimated cost of $16 M to $17 M).

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26.2.2 Environmental and Permitting

· Additional baseline data collection is on going in support of the EA for the proposed mine plan and infrastructure as outlined in the NGD 2012 PEA, and should include areas such as hydrological, air quality, meteorology, aquatics, vegetation and fauna studies, ecosystem mapping, soil surveys, and archaeological surveys. Studies will be required in support of the Project EA, including acid rock drainage and metals leaching studies and water quality modelling (estimated total cost of $2 M to $3 M).

· Environmental studies and permitting includes studies required to complete an Environmental Impact Statement as per provincial and federal environmental assessment acts will be completed in 2013.  This includes detailed environmental, social and economic analysis culminating in completion of baseline reports and an EIS in late 2013 (estimated cost of $10 M to $11 M).

26.2.3 Government and Community

· Community and Government engagement will include maintenance and staffing of a local community office as well as support for community development projects and local community engagement and communications.  This includes First Nations engagement that will include capacity funding for local aboriginal peoples with an interest in the project as per Exploration Cooperation Agreements and additional capacity funding as part of the Environmental Assessment process and exploration program (estimated cost of $6 M to $7 M).

26.2.4 Project Development Planning

· Additional studies will be required to advance the planning for the project to prepare for implementation (estimated cost of $4M to $5M).

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27.0 REFERENCES

27.1 References

A.Giblett and J.Seidel, Measuring, Predicting and Managing Grinding Media Wear,

               SAG 2011, Vancouver, BC, Canada, September 2011

Adams, J. and Halchuk, S., 2003.  Fourth generation seismic hazard maps of Canada:

               Values for over 650 Canadian localities intended for the 2005 National Building

               Code of Canada, Geological Survey of Canada, Open File 4459.

Allnorth Consultants Ltd., Blackwater Project North Haul Route Analysis, April 18,

               2012

AMEC, New Gold Inc. Blackwater Process Trade-off Study, January 24, 2012

BC Hydro,

               http://www.bchydro.com/etc/medialib/internet/documents/appcontent/your_acc

               ount/BC_Hydro_Electric_Tariff.Par.0001.File.BC-Hydro-Tariff.pdf

               Accessed May 11, 2012

Canadian Institute of Mining, Metallurgy and Petroleum (CIM), 2003:  Estimation of

               Mineral Resources and Mineral Reserves, Best Practice Guidelines:  Canadian

               Institute of Mining, Metallurgy and Petroleum, November 23, 2003,

               http://www.cim.org/committees/estimation2003.pdf.

Canadian Institute of Mining, Metallurgy and Petroleum (CIM), 2010:  CIM Standards

               for Mineral Resources and Mineral Reserves, Definitions and Guidelines:  

               Canadian Institute of Mining, Metallurgy and Petroleum, November 2010,

               http://www.cim.org/UserFiles/File/CIM_DEFINITON_STANDARDS_Nov_2010.pdf

Canadian Securities Administrators (CSA), 2011: National Instrument 43-101,

               Standards of Disclosure for Mineral Projects, Canadian Securities

               Administrators.

Diakow, L.J. and V.M. Levson (1997):  Bedrock and Surfical Geology of the Southern

               Nechako Plateau, Central British Columbia (93F/2,3,6,7), Geoscience Map

               1997-2. (Map)

Diakow, L.J., and Webster, I.C.L. (1994):  Geology of the Fawnie Creek Map Area

               (93F/3); in Geological Fieldwork 1993, Grant, B and Newell, J.M., Editors, B.C.

               Ministry of Energy Mines, and Petroleum Resources, Paper 1994-1, pages 15-26.

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Diakow, L.J., Webster, I.C.L., Richards, T.A. and Tipper, H.W. (1997):  Geology of the

               Fawnie and Nechako Ranges, Southern Nechako Plateau, Central British

               Columbia (96F/2,3,6,7); in Interior Plateau Geoscience Project:  Summary of

               Geological, Geochemical and Geophysical Studies, Newell, J.M. and Diakow

               L.G., Editors.  B.C. Ministry of Employment and Investment, Paper 1997-2.

FLSmidth Dawson Metallurgical Laboratories., P-4257 Final Report on Metallurgical

               Test Work on Sulphide, Oxide and Transition Composite Samples from New

               Gold’s Blackwater Property in Central British Columbia, April 17, 2012

G&T Metallurgical Services Ltd., KM2885 Preliminary Metallurgical Testing Blackwater

               Project, May 11, 2011

G&T Metallurgical Services Ltd., KM3071 Cyanidation Bottle Roll Testing Blackwater

               Project, September 12, 2011

G&T Metallurgical Services Ltd., KM3145 Ore Hardness Testing Blackwater Project,

               September 26, 2011

G&T Metallurgical Services Ltd., KM3156 Alternative Flowsheet Evaluation Blackwater

               Project Samples, May 15, 2012

Inspectorate Metallurgical Division, 1001302 Preliminary Metallurgical Testing of

               Samples from the Richfield Ventures Corp., Blackwater Project of Central

               British Columbia, February 25, 2011

Knight Piésold, KP Transmission study letter report, 2011 Richfield study.

Knight Piésold Ltd, January 2011, Blackwater Gold Project – Tailings Storage Facility

               Site Alternative Assessment (VA10-00235)

Levson, V. M., Weary, G. F., & Giles, T. R. (1995). Surficial Geology and Drift

               Exploration Studies in the Tsacha Lake and Chedakuz Creek Areas. BCGS

               Geological Fieldwork.

Mate, D. J., & Levson, V. M. (2000). Quaternary stratigraphy and history of the Ootsa

               Lake-Cheslatte River area, Nechako Plateau, central British Columbia. Can. J.

               Earth Sci

McClelland Laboratories., MLI3597 Interim Results, 2012

Plouffe, A., & Levson, V. M. (2000). Late Quaternary glacial and interglacial

               environments of the Nechako River. Can. J. Earth Sci.

SGS Canada., 50179-001 Interim Results, 2012

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Sillitoe, R.H., Hedenquist, J.W. (2003):  Linkages between volcanotectonic settings,

               ore-fluid compositions, and epithermal precious metal deposits:  Society of

               Economic Geologists Special Publication 10, p. 315-343.

27.2 Glossary

27.2.1 Abbreviations

Abbreviation	Definition
®	registered name
AA	atomic absorption spectroscopy
ABA	acid base accounting
AOI	area of interest
ARD	acid rock drainage
AuEq	gold equivalent grade
BCEAO	BC Environmental Assessment Office
BCWQO	B.C. Water Quality Objectives
BGC	biogeoclimatic
CCME	Canadian Council of Ministries of Environment
CIM	Canadian Institute of Mining, Metallurgy and Petroleum
CIP	carbon in pulp
CM	construction management
COS	coarse ore stockpile
COSEWIC	Committee on the Status of Endangered Wildlife in Canada
CRM	certified reference material
CTF	cross training factor
DCF	discounted cash flow
EA	Environmental Assessment
EAC	Environmental Assessment Certificate Application
ECD	environmental control dam
EDA	exploratory data analysis
EPCM	engineering procurement construction management
FA	fire assay
FCL	flotation concentrate leach
FS	Feasibility Study
FSR	forest service road
GPS	global positioning system
HL	heap leach
ICC	International Cyanide Management Code

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Abbreviation	Definition
ICMI	International Cyanide Management Institute
ICP	inductively-couple plasma
ICP-AES	inductively-coupled plasma atomic emission spectrometry
ICP-MS	inductively-coupled plasma mass spectrometry
IMC	Independent Mining Consultants
IP	induced polarization
IRR	internal rate of return
ISQG	Interim Sediment Quality Guideline
KP	Knight Piésold
LEL	lowest effect level
L–G	Lerchs–Grossmann
LiDAR	aerial light detection and ranging
LOM	life of mine
LSA	local study area
ML	metal leaching
MMER	Metal Mining Effluent Regulations
MTO	Mineral Titles Online
MYAB	multi year area based
NaCN	sodium cyanide
NAG	non-acid generating
NI 43-101	Canadian National Instrument 43-101 “Standards of Disclosure for Mineral Projects”
NPV5	net present value discounted at 5% (pre-tax)
NSR	net smelter return
P.Eng. or P.E.	Professional Engineer
P.Geol or P.Geo	Professional Geologist
PAG	potentially acid generating
PEA	Preliminary Economic Assessment
PEL	permissible exposure level
PMA	particle mineralogical analysis
QA/QC	quality assurance/quality control
QEMSCAN	Quantitative Evaluation of Minerals by Scanning Electron Microscopy
RC	reverse circulation
ROM	run of mine
RQD	rock quality designation
RSA	regional study area
SAG	semi-autogenous grinding
SARA	Species at Risk Act
TK/TLU	traditional knowledge and traditional land use

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Abbreviation	Definition
TSF	tailings storage facility
UCS	unconfined compressive strength
WAD	weak acid dissociable cyanide
WBS	work breakdown structure
WOL	whole ore leach
XRF	X-ray fluorescence

27.2.2 Glossary Table

Term	Definition
acid rock drainage/ acid mine drainage	Characterized by low pH, high sulphate, and high iron and other metal species.
adjacent property	A property in which the issuer does not have an interest; has a boundary reasonably proximate to the property being reported on; and has geological characteristics similar to those of the property being reported on
advanced argillic alteration	Consists of kaolinite + quartz + hematite + limonite.  feldspars leached and altered to sericite.  The presence of this assemblage suggests low pH (highly acidic) conditions.  At higher temperatures, the mineral pyrophyllite (white mica) forms in place of kaolinite
advanced property	A means a property that has mineral reserves, or mineral resources the potential economic viability of which is supported by a preliminary economic assessment, a pre-feasibility study or a feasibility study.
alluvium	Unconsolidated terrestrial sediment composed of sorted or unsorted sand, gravel, and clay that has been deposited by water.
ANFO	A free-running explosive used in mine blasting made of 94% prilled aluminum nitrate and 6% No. 3 fuel oil.
aquifer	A geologic formation capable of transmitting significant quantities of groundwater under normal hydraulic gradients.
argillic alteration	Introduces any one of a wide variety of clay minerals, including kaolinite, smectite and illite.  Argillic alteration is generally a low temperature event, and some may occur in atmospheric conditions
autogenous grinding	The process of grinding in a rotating mill which uses as a grinding medium large pieces or pebbles of the ore being ground, instead of conventional steel balls or rods.
azimuth	The direction of one object from another, usually expressed as an angle in degrees relative to true north.  Azimuths are usually measured in the clockwise direction, thus an azimuth of 90 degrees indicates that the second object is due east of the first.
background concentration	Naturally-occurring concentrations of compounds of environmental concern

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Term	Definition
ball mill	A piece of milling equipment used to grind ore into small particles.  It is a cylindrical shaped steel container filled with steel balls into which crushed ore is fed.  The ball mill is rotated causing the balls themselves to cascade, which in turn grinds the ore.
beneficiation	Physical treatment of crude ore to improve its quality for some specific purpose. Also called mineral processing.
bullion	Unrefined gold and/or silver mixtures that have been melted and cast into a bar or ingot.
carbon-in-pulp (CIP)	A method of recovering gold and silver from fine ground ore by adsorption of the precious metals onto fine carbon in an agitated tank of ore solids/solution slurry. This recovery step in the process follows the leaching process which is done in similarly agitated tanks, but without contained carbon.
comminution/crushing/grinding	Crushing and/or grinding of ore by impact and abrasion. Usually, the word "crushing" is used for dry methods and "grinding" for wet methods. Also, "crushing" usually denotes reducing the size of coarse rock while "grinding" usually refers to the reduction of the fine sizes.
concentrate	The concentrate is the valuable product from mineral processing, as opposed to the tailing, which contains the waste minerals. The concentrate represents a smaller volume than the original ore
cut-off grade	A grade level below which the material is not “ore” and considered to be uneconomical to mine and process. The minimum grade of ore used to establish reserves.
cyanidation	A method of extracting gold or silver by dissolving it in a weak solution of sodium cyanide.
data verification	The process of confirming that data has been generated with proper procedures, has been accurately transcribed from the original source and is suitable to be used for mineral resource and mineral reserve estimation
density	The mass per unit volume of a substance, commonly expressed in grams/ cubic centimeter.
depletion	The decrease in quantity of ore in a deposit or property resulting from extraction or production.
development property	a property that is being prepared for mineral production or a material expansion of current production, and for which economic viability has been demonstrated by a pre-feasibility or feasibility study.
dilution	Waste of low-grade rock which is unavoidably removed along with the ore in the mining process.
disclosure	Any oral statement or written disclosure made by or on behalf of an issuer and intended to be, or reasonably likely to be, made available to the public in a jurisdiction of Canada, whether or not filed under securities legislation, but does not include written disclosure that is made available to the public only by reason of having been filed with a government or agency of government pursuant to a requirement of law other than securities legislation.

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Term	Definition
discounted cash flow (DCF)	Concept of relating future cash inflows and outflows over the life of a project or operation to a common base value thereby allowing more validity to comparison of projects with different durations and rates of cash flow.
early-stage exploration property	A property for which the technical report being filed has no current mineral resources or mineral reserves defined; and no drilling or trenching proposed
easement	Areas of land owned by the property owner, but in which other parties, such as utility companies, may have limited rights granted for a specific purpose.
effective date	With reference to a technical report, the date of the most recent scientific or technical information included in the technical report.
electrowinning.	The removal of precious metals from solution by the passage of current through an electrowinning cell.  A direct current supply is connected to the anode and cathode.  As current passes through the cell, metal is deposited on the cathode.  When sufficient metal has been deposited on the cathode, it is removed from the cell and the sludge rinsed off the plate and dried for further treatment.
elution	Recovery of the gold from the activated carbon into solution before zinc precipitation or electro-winning.
EM	Geophysical method, electromagnetic system, measures the earth's response to electromagnetic signals transmitted by an induction coil
encumbrance	An interest or partial right in real property which diminished the value of ownership, but does not prevent the transfer of ownership.  Mortgages, taxes and judgements are encumbrances known as liens.  Restrictions, easements, and reservations are also encumbrances, although not liens.
exploration information	Geological, geophysical, geochemical, sampling, drilling, trenching, analytical testing, assaying, mineralogical, metallurgical, and other similar information concerning a particular property that is derived from activities undertaken to locate, investigate, define, or delineate a mineral prospect or mineral deposit
feasibility study	A comprehensive study of a mineral deposit in which all geological, engineering, legal, operating, economic, social, environmental, and other relevant factors are considered in sufficient detail that it could reasonably serve as the basis for a final decision by a financial institution to finance the development of the deposit for mineral production.
flotation	Separation of minerals based on the interfacial chemistry of the mineral particles in solution. Reagents are added to the ore slurry to render the surface of selected minerals hydrophobic. Air bubbles are introduced to which the hydrophobic minerals attach. The selected minerals are levitated to the top of the flotation machine by their attachment to the bubbles and into a froth product, called the "flotation concentrate." If this froth carries more than one mineral as a designated main constituent, it is called a "bulk float". If it is selective to one constituent of the ore, where more than one will be floated, it is a "differential" float.

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Term	Definition
floating cone	A method that searches through the block model for all blocks and assesses the value of the inverted kind that have to be mined to expose them.  If the value of a kind is positive, it is mined out and all the blocks it includes a changed to their block.  The search then continues.
flowsheet	The sequence of operations, step by step, by which ore is treated in a milling, concentration, or smelting process.
footwall	The wall or rock on the underside of a vein or ore structure.
free milling	Ores of gold or silver from which the precious metals can be recovered by concentrating methods without resort to roasting or chemical treatment.
frother	A type of flotation reagent which, when dissolved in water, imparts to it the ability to form a stable froth
gangue	The fraction of ore rejected as tailing in a separating process.  It is usually the valueless portion, but may have some secondary commercial use
gravity separation	Exploitation of differences in the densities of particles to achieve separation. Machines utilizing gravity separation include jigs and shaking tables.
hanging wall	The wall or rock on the upper or top side of a vein or ore deposit.
heap leaching	A process whereby valuable metals, usually gold and silver, are leached from a heap or pad of crushed ore by leaching solutions percolating down through the heap and collected from a sloping, impermeable liner below the pad.
Indicated Mineral Resource	An ‘Indicated Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit.  The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough for geological and grade continuity to be reasonably assumed.
Inferred Mineral Resource	An ‘Inferred Mineral Resource’ is that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity.  The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes.
internal rate of return (IRR)	The rate of return at which the Net Present Value of a project is zero; the rate at which the present value of cash inflows is equal to the present value of the cash outflows.
IP	Geophysical method, induced polarization; used to directly detect scattered primary sulfide mineralization.  Most metal sulfides produce IP effects, e.g. chalcopyrite, bornite, chalcocite, pyrite, pyrrhotite

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Term	Definition
Knelson concentrator	a high-speed centrifuge that combines centrifugally enhanced gravitational force with a patented fluidization process to recover precious metals
liberation	Freeing, by comminution, of particles of specific mineral from their interlock with other constituents of the ore.
life of mine (LOM)	Number of years that the operation is planning to mine and treat ore, and is taken from the current mine plan based on the current evaluation of ore reserves.
Measured Mineral Resource	A ‘Measured Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, and physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough to confirm both geological and grade continuity.
Merrill-Crowe (M-C) circuit	A process which recovers precious metals from solution by first clarifying the solution, then removing the air contained in the clarified solution, and then precipitating the gold and silver from the solution by injecting zinc dust into the solution. The valuable sludge is collected in a filter press for drying and further treatment
mill	Includes any ore mill, sampling works, concentration, and any crushing, grinding, or screening plant used at, and in connection with, an excavation or mine.
mineral project	Any exploration, development or production activity, including a royalty or similar interest in these activities, in respect of diamonds, natural solid inorganic material, or natural solid fossilized organic material including base and precious metals, coal, and industrial minerals
Mineral Reserve	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.
Mineral Resource	A Mineral Resource is 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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Term	Definition
mining claim	A description by boundaries of real property in which metal ore and/or minerals may be located.
net present value (NPV)	The present value of the difference between the future cash flows associated with a project and the investment required for acquiring the project.  Aggregate of future net cash flows discounted back to a common base date, usually the present.  NPV is an indicator of how much value an investment or project adds to a company.
net smelter return royalty (NSR)	A defined percentage of the gross revenue from a resource extraction operation, less a proportionate share of transportation, insurance, and processing costs.
open pit	A mine that is entirely on the surface. Also referred to as open-cut or open-cast mine.
ounce (oz) (troy)	Used in imperial statistics.  A kilogram is equal to 32.1507 ounces.  A troy ounce is equal to 31.1035 grams.
overburden	Material of any nature, consolidated or unconsolidated, that overlies a deposit of ore that is to be mined.
petrography	Branch of geology that deals with the description and classification of rocks.
plant	A group of buildings, and especially to their contained equipment , in which a process or function is carried out; on a mine it will include warehouses, hoisting equipment, compressors, repair shops, offices, mill or concentrator.
preliminary economic assessment	A study, other than a pre-feasibility or feasibility study, that includes an economic analysis of the potential viability of mineral resources
preliminary feasibility study, pre-feasibility study	A comprehensive study of the viability of a mineral project that has advanced to a stage where the mining method, in the case of underground mining, or the pit configuration, in the case of an open pit, has been established and an effective method of mineral processing has been determined, and includes a financial analysis based on reasonable assumptions of technical, engineering, legal, operating, economic, social, and environmental factors and the evaluation of other relevant factors which are sufficient for a qualified person, acting reasonably, to determine if all or part of the mineral resource may be classified as a mineral reserve
pyrometallurgy	A type of extractive metallurgy where furnace treatments at high temperature are used to separate the metal values from an ore or concentrate.  The waste product is removed as slag and/or gases.  Smelting and refining are common pyrometallurgical processes.
reclamation	The restoration of a site after mining or exploration activity is completed.
refining	A high temperature process in which impure metal is reacted with flux to reduce the impurities.  The metal is collected in a molten layer and the impurities in a slag layer.  Refining results in the production of a marketable material.

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Term	Definition
refractory	Gold mineralization normally requiring more sophisticated processing technology for extraction, such as roasting or autoclaving under pressure.
Resistivity	Observation of electric fields caused by current introduced into the ground as a means of studying earth resistivity in geophysical exploration. Resistivity is the property of a material that resists the flow of electrical current
right-of-way	A parcel of land granted by deed or easement for construction and maintenance according to a designated use.  This may include highways, streets, canals, ditches, or other uses
royalty	An amount of money paid at regular intervals by the lessee or operator of an exploration or mining property to the owner of the ground. Generally based on a specific amount per tonne or a percentage of the total production or profits. Also, the fee paid for the right to use a patented process.
run-of-mine	A term used to describe ore of average grade for the deposit.
semi-autogenous grinding (SAG)	A method of grinding rock into fine powder whereby the grinding media consists of larger chunks of rocks and steel balls.
specific gravity	The weight of a substance compared with the weight of an equal volume of pure water at 4°C.
strike length	The horizontal distance along the long axis of a structural surface, rock unit, mineral deposit or geochemical anomaly.
strip ratio	The ratio of waste tons to ore tons mined calculated as total tonnes mined less ore tonnes mined divided by ore tonnes mined.
tailings	Material rejected from a mill after the recoverable valuable minerals have been extracted.
World Geodetic Reference System of 1984 (WGS-84)-	The United States Defense Mapping Agency's Datum. This datum is a global datum based on electronic technology which is still to some degree classified. Data on the relationship of as many as 65 different datums to WGS-84 is available to the public. As a result, WGS-84 is becoming the base datum for the processing and conversion of data from one datum to any other datum. The Global Positioning System (GPS) is based on this datum.
written disclosure	Any writing, picture, map, or other printed representation whether produced, stored or disseminated on paper or electronically, including websites

27.2.3 Chemical Symbols Table

Symbol	Element or Compound
Ag	silver
Al	aluminum
As	arsenic
Au	gold

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Symbol	Element or Compound
B	boron
Ba	barium
Be	beryllium
Bi	bismuth
C	carbon
Ca	calcium
CaCO3	calcium carbonate
CaO	calcium oxide
Cd	cadmium
Ce	cerium
Cl	chlorine
CN	cyanide
Co	cobalt
Cr	chromium
Cs	cesium
Cu	copper
Fe	iron
FeOx	iron oxides
Ga	gallium
Ge	germanium
H	hydrogen
Hf	hafnium
Hg	mercury
In	indium
K	potassium
La	lanthium
Li	lithium
Mg	magnesium
Mn	manganese
Mn(OH)2	manganous hydroxide
MnO2	manganese dioxide
Mo	molybdenum
N	nitrogen
Na	sodium
Nb	niobium
NH3	ammonia
Ni	nickel
NOx	nitrogen oxide compounds

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Symbol	Element or Compound
O2	oxygen
P	phosphorus
Pb	lead
Pd	palladium
Pt	platinum
Rb	rubidium
Re	rhenium
S	sulfur
Sb	antimony
Sc	scandium
Se	selenium
Sn	tin
SO2	sulfur dioxide
Sr	strontium
Ta	tantalum
Te	tellurium
Th	thorium
Ti	titanium
Tl	thallium
U	uranium
V	vanadium
W	tungsten
Y	yttrium
Zn	zinc
Zr	zirconium

27.2.4 Metric Measurements Table

Measurement	Definition
'	seconds (geographic)
'	foot/feet
"	minutes (geographic)
"	inches
#	number
%	percent
/	per
<	less than

Project No.:  169229
October 2012

Page 27-13

New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Measurement	Definition
>	greater than
µm	micrometer (micron)
a	annum/ year
Å	angstroms
asl	above sea level
BQ	36.5 mm size core
c.	circa
d	day
d/wk	days per week
dmt	dry metric tonne
fineness	parts per thousand of gold in an alloy
g	gram
g/cm3	grams per cubic centimeter
g/m3	grams per cubic meter
Ga	billion years ago
ha	hectares
HP	horsepower
HQ	63.5 mm size core
kg/m3	kilograms per cubic meter
km	kilometer
km2	square kilometers
koz	thousand ounces
kV	kilovolt
kVA	kilovolt–ampere
kW	kilowatt
kWh	kilowatt hour
lb	pound
M	million
m	meter
m3	cubic meter
m3/hr	cubic meters per hour
Ma	million years ago
mesh	size based on the number of openings in one inch of screen
mi	mile/miles
Mlb	million pounds
Mm	million meters
mm	millimeter/millimeters

Project No.:  169229
October 2012

Page 27-14

New Gold Inc. Blackwater Project British Columbia, Canada NI 43-101 Technical Report on Preliminary Economic Assessment

Measurement	Definition
Moz	million ounces
Mt	million tonnes
Mt/a	million tonnes per annum
MW	megawatts
NQ	47.6 mm size core
º	degrees
ºC	degrees Celsius
oz	ounce/ounces (troy ounce)
oz/t	ounces per tonne
p	passing
pH	measure of the acidity or alkalinity of a solution
pop	population
ppb	parts per billion
ppm	parts per million
PQ	85 mm size core
t	metric tonne
t/a	tonnes per annum (tonnes per year)
t/d	tonnes per day
t/h	tonnes per hour
t/m3	tonnes per cubic meter
TDS	total dissolved solids
TSS	total suspended solids
wt%	weight percent

Project No.:  169229
October 2012

Page 27-15