newaftontechnical-2010.htm

EX-99.1 2 newaftontechnical-2010.htm NEW AFTON TECHNICAL REPORT OF DECMBER 31, 2010 newaftontechnical-2010.htm

Exhibit 99.1

SCOTT WILSON RPA WWW.SCOTTWILSON.COM

TABLE OF CONTENTS

PAGE

1	SUMMARY	1-1
	Executive Summary	1-1
	Technical Summary	1-9
2	INTRODUCTION	2-1
3	RELIANCE ON OTHER EXPERTS	3-1
4	PROPERTY DESCRIPTION AND LOCATION	4-1
5	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY	5-1
6	HISTORY	6-1
7	GEOLOGICAL SETTING	7-1
	Regional Geology	7-1
	Property Geology	7-2
8	DEPOSIT TYPES	8-1
9	MINERALIZATION	9-1
10	EXPLORATION	10-1
11	DRILLING	11-1
12	SAMPLING METHOD AND APPROACH	12-1
13	SAMPLE PREPARATION, ANALYSES AND SECURITY	13-1
14	DATA VERIFICATION	14-1
15	ADJACENT PROPERTIES	15-1
16	MINERAL PROCESSING AND METALLURGICAL TESTING	16-1
	Metallurgical Testing	16-1
	Mineral Processing	16-15
17	MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES	17-1
	Mineral Resources	17-1
	Mineral Reserves	17-19
18	OTHER RELEVANT DATA AND INFORMATION	18-1
19	ADDITIONAL REQUIREMENTS	19-1
	Mining Operations	19-2
	Mineral Processing	19-40
	Life of Mine Plan	19-40
	Markets	19-43
	Contracts	19-45
	Environmental Considerations	19-45

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Technical Report NI 43-101 – December 31, 2009

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	Taxes	19-53
	Capital and Operating Cost Estimates	19-54
	Project Execution Plan	19-62
	Economic Analysis	19-62
20	INTERPRETATION AND CONCLUSIONS	20-1
21	RECOMMENDATIONS	21-1
22	REFERENCES	22-1
23	DATE AND SIGNATURE PAGE	23-1
24	CERTIFICATE OF QUALIFIED PERSON	24-1
25	APPENDIX 1	25-1

LIST OF TABLES

PAGE

Table 1-1	After-Tax Cash Flow	1-5
Table 1-2	Sensitivity Analyses	1-8
Table 1-3	Mineral Resource Summary – September 2009	1-13
Table 1-4	Mineral Resource Summary – 0.4% CuEq Cut-Off	1-14
Table 1-5	Mineral Reserve Estimate – September 2009	1-15
Table 4-1	Mineral Tenures	4-2
Table 11-1	Drill Summary by Campaign	11-1
Table 15-1	Ajax Deposit – Mineral Resource Estimate	15-1
Table 16-1	Percentage of Ore Type	16-1
Table 16-2	Bulk Samples	16-3
Table 16-3	Grinding Index Summary	16-5
Table 16-4	Locked Cycle Test Summary	16-7
Table 16-5	Concentrate Assays	16-7
Table 16-6	Variability (Batch) Test Summary Results	16-10
Table 17-1	Mineral Resource Summary	17-2
Table 17-2	Block Model Geometry	17-4
Table 17-3	Statistical Summary of Sample Assay Data	17-5
Table 17-4	Top Cuts	17-9
Table 17-5	Variogram Models	17-10
Table 17-6	Bulk Density	17-12
Table 17-7	Mineral Resource Estimate Comparison	17-14
Table 17-8	Mineral Resource Summary (CuEq Cut-Off)	17-15
Table 17-9	2009 vs. 2006 Estimate Comparison	17-18
Table 17-10	Mineral Reserve Estimate – September 2009	17-19
Table 17-11	Drawpoint Production by Resource Classification	17-24
Table 17-12	Tonnage and Grade of Material from Mine Construction Activities	17-25
Table 17-13	Mineral Reserve Estimate	17-26
Table 19-1	LOM Ore and Waste Tonnage Summary	19-28
Table 19-2	Underground Mobile Equipment	19-38
Table 19-3	Capital Equipment for Production	19-39

New Gold Inc. – New Afton Project

Technical Report NI 43-101 – December 31, 2009

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Table 19-4	Sustaining Capital U/G Equipment	19-39
Table 19-5	LOMP Development	19-40
Table 19-6	LOMP development Tonnage	19-41
Table 19-7	LOM Mill Feed	19-42
Table 19-8	LOM Metal Recovery	19-42
Table 19-9	Concentrate Production	19-43
Table 19-10	Metal Price and Exchange Rate Assumptions	19-44
Table 19-11	Authorizations and Licences	19-47
Table 19-12	Closure Cost Estimate	19-53
Table 19-13	Construction Capital Estimate	19-55
Table 19-14	Site Development Capital Estimate	19-55
Table 19-15	Process Capital Estimate	19-56
Table 19-16	Electrical Capital Estimate	19-56
Table 19-17	Tailings Capital Estimate	19-57
Table 19-18	Surface Services Capital Estimate	19-57
Table 19-19	Mining Capital Estimate	19-58
Table 19-20	Indirect Capital Estimate	19-58
Table 19-21	Sustaining Capital by Area	19-59
Table 19-22	Annual Sustaining Capital	19-60
Table 19-23	Annual Sustaining Capital Breakdown	19-60
Table 19-24	Operating Costs	19-61
Table 19-25	Manpower	19-61
Table 19-26	After-Tax Cash Flow	19-64
Table 19-27	Sensitivity Analyses	19-67

LIST OF FIGURES

PAGE

Figure 1-1	Sensitivity Analysis	1-8
Figure 4-1	Location Map	4-4
Figure 4-2	Mineral Titles	4-5
Figure 7-1	Regional Geology	7-3
Figure 7-2	Afton Pit Geology	7-4
Figure 7-3	Property Geology	7-5
Figure 11-1	Drill Hole Locations and Resource Outline	11-6
Figure 16-1	Process Flowheet	16-18
Figure 19-1	Location of Mine Workings as Related to Surface	19-4
Figure 19-2	Location of the Mining Blocks	19-5
Figure 19-3	Block Undercut Outlines	19-10
Figure 19-4	Caving Stability Chart	19-11
Figure 19-5	Surface Subsidence Projections	19-15
Figure 19-6	Subsidence Limits Based on Empirical Assessment Methods	19-16
Figure 19-7	Undercut Design	19-22
Figure 19-8	Drawpoint Arrangement	19-25
Figure 19-9	Typical Drawpoint Support	19-26
Figure 19-10	Ore and Waste Tonnage	19-28
Figure 19-11	Conveyor Gallery Layout	19-30

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Figure 19-12	Conveyor Decline Cross-Section	19-31
Figure 19-13	Sensitivity Analysis	19-67

LIST OF APPENDIX FIGURES & TABLES

PAGE

Figure A1-1

Composite Statistics, Histograms and Probability Plots

25-2

New Gold Inc. – New Afton Project Technical Report NI 43-101 – December 31, 2009	Page iv

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1 SUMMARY

EXECUTIVE SUMMARY

Scott Wilson Roscoe Postle Associates Inc. (Scott Wilson RPA) was retained by Mr. Mark Petersen, Vice President, Exploration, of New Gold Inc. (New Gold), to prepare an independent Technical Report and Feasibility Update on the New Afton Cu-Au-Ag Project (the Project) near Kamloops, British Columbia, Canada. The purpose of this report is to support the public disclosure of updated Mineral Resources and Mineral Reserves for the Project. This Technical Report conforms to NI 43-101 Standards of Disclosure for Mineral Projects (NI 43-101). Scott Wilson RPA visited the property on September 3 and 4, 2009.

New Afton is a former producer which operated an open pit mine from 1977 to 1987. New Gold acquired an option on the New Afton property in 1999 and began exploration work in 2000. The Project is in the development stage and is a planned 11,000 tpd (nominal) block cave operation scheduled to commence milling in mid-2012. Copper and gold in concentrates will be produced at the Project. Development commenced in 2007 but, except for engineering and other limited activities, was halted due to the global economic conditions in the fourth quarter of 2008. Underground development recommenced in the first quarter of 2009 and was underway at the time of the site visit. Surface construction is scheduled to restart in 2011.

CONCLUSIONS

Scott Wilson RPA has carried out an audit of the Mineral Resource and Mineral Reserve estimates for the New Afton Project, and makes the following conclusions:

·	The drilling has been conducted in a manner consistent with industry standards.

·	The core handling, logging and sampling have been carried out in a manner consistent with industry standards.

·	Sampling and assaying has been conducted in a manner appropriate for the deposit type and mineralization, and using industry-standard practices.

·	Assay quality assurance/quality control (QA/QC) protocols are consistent with common industry standards, and no evidence of persistent or systematic errors has been found. However, the QA/QC has been inconsistently monitored and requires more timely review.

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·	Density measurements are made in sufficient numbers and with an appropriate methodology.

·	The database is generally free of serious errors, although several inconsistencies have been found. At the time of writing, a new database system was being installed, which will implement more rigorous validation procedures.

·	The Mineral Resource estimate has been carried out using generally reasonable assumptions and parameters, and accepted practices.

Geological interpretation has resulted in improvements to the understanding of several features of the deposit, particularly the faulting and the De-mineralized Zone (DMZ). However, interpretations of the mineralized zone boundaries have not been maintained, resulting in the implementation of a grade shell approach to constraining the grade interpolation. In Scott Wilson RPA’s opinion, it would be preferable to base the resource boundaries on lithological and mineralogical characteristics, if possible.

·	The updated Mineral Resource estimate is not materially different from the 2006 estimate used in the Feasibility Study completed by Hatch Ltd. in 2007, and should not result in a major change to the Mineral Reserves.

·	The assay database for deleterious elements has been markedly improved since the last estimate but still lags behind the economic elements in terms of numbers of samples and coverage. Improvements could still be made to the block grade estimates for deleterious components.

·	The Mineral Resources should be quoted at a CuEq cut-off rather than a Cu cut-off.

·	The Mineral Reserve is estimated to 47.4 million tonnes of Probable Mineral Reserves grading 0.95% Cu, 0.69 g/t Au, and 2.03 g/t Ag.

·	The Mineral Reserves are considered to be amenable to underground mining through the use of block caving at a production rate of 4 Mtpa of ore.

·	The preproduction capital cost is estimated to be C$460.9 million plus sustaining capital of $172.3 million and not including sunk expenditures.

·	The estimated Life of Mine (LOM) operating cost is C$17.27 per tonne milled.

·	Considering the Project on a stand-alone basis, the undiscounted after-tax cash flow totals C$707 million over the mine life, and simple payback occurs approximately 4.5 years after the start of production.

The Total Cash Cost is US$0.84 per pound of copper. The mine life capital unit cost is US$0.69 per pound of copper and the byproduct credits are US$0.89 per pound of copper, for a Total Production Cost of US$0.64 per pound of copper. Average annual production during operation is 72 million pounds of copper and 76,000 ounces of gold.

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·	The after-tax net present value (NPV) at a 5% discount rate is C$374 million, and the internal rate of return (IRR) is 19%.

RECOMMENDATIONS

Scott Wilson RPA makes the following recommendations:

·	Based upon the assumptions and estimates within this study, the continued development of the Project is recommended.

·	Database validation protocols should be enhanced and rigorously observed.

·	Geological interpretation and wireframing should continue with focus on the zones of mineralization, in order to improve the constraints to the grade interpolations. If possible, the grade shell should be replaced by wireframes based on geological characteristics.

·	The Mineral Resource estimate should be quoted at a 0.4% CuEq cut-off. Changing the cut-off does not materially affect the Mineral Reserves.

·	Analytical work on the deleterious elements should continue in order to allow improvements to be made to the block models for these components.

·	The review of the project execution plan which is currently underway should be continued and completed including incorporation of all of the modifications since completion of the 2007 Feasibility Study.

ECONOMIC ANALYSIS

A Cash Flow Projection has been generated from the LOM production schedule and capital and operating cost estimates, and is summarized in Table 1-1. A summary of the key criteria is provided below.

ECONOMIC CRITERIA

Revenue

·	4 Mpta mining from underground

·	Mill recovery by zone, as indicated by testwork, averaging 87.8%.

·	Exchange rate US$1.00 = C$0.85.

·	Metal prices as shown in Table 19-10.

·	Net Smelter Return includes smelting, refining, transport, and penalty treatment costs.

·	Revenue is recognized at the time of mill production

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Costs

·	Pre-production period: 30 months (January 2010 to July 2012).

·	Mine life: 12 years.

·	LOM production plan as summarized in Tables 19-5 to 19-7.

·	Pre-production capital C$460.9 million.

·	Average operating cost over the mine life is C$17.27 per tonne milled.

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New Gold Inc. – New Afton Project	Page 1-5
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CASH FLOW ANALYSIS

Considering the Project on a stand-alone basis, the undiscounted after-tax cash flow totals C$707 million over the mine life, and simple payback occurs after approximately 4.5 years.

The after tax NPV at a 5% discount rate is C$374 million, and the IRR is 19%.

SENSITIVITY ANALYSIS

Project risks can be identified in both economic and non-economic terms. Key economic risks were examined by running cash flow sensitivities:

·	Metal prices

·	Operating costs (Total Cash Cost)

·	Capital costs

The after-tax 5% NPV sensitivity over the base case has been calculated for -20% to +20% variations. The sensitivities are shown in Figure 1-1 and Table 1-2. The Project is most sensitive to metal price and less sensitive to capital and operating costs, which are almost equally sensitive.

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Technical Report NI 43-101 – December 31, 2009

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Figure 1-1 Sensitivity Analysis

TABLE 1-2 SENSITIVITY ANALYSES

New Gold Inc. – New Afton Project

Sensitivity to Metal Price
Cu US$/lb	NPV @ 0% millions	NPV @ 5% millions	NPV @ 8% millions
1.64	302	104	28
1.85	505	239	136
2.05	707	374	244
2.26	910	510	353
2.46	1,112	644	460

Sensitivity to Operating Cost
$/tonne	NPV @ 0% millions	NPV @ 5% millions	NPV @ 8% millions
13.81	825	452	306
15.54	766	413	275
17.27	707	374	244
18.99	648	336	213
20.72	591	298	184

Sensitivity to Capital Cost
$ Millions	NPV @ 0% millions	NPV @ 5% millions	NPV @ 8% millions
513	835	484	345
577	771	429	295
641	707	374	244
705	645	321	195
769	581	266	145

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TECHNICAL SUMMARY

PROPERTY DESCRIPTION AND LOCATION

The New Afton Project is located at Latitude 50°39’ north and Longitude 120°31’ west, approximately 350 km northeast of Vancouver and 10 km west of the City of Kamloops, in the South-Central Interior of British Columbia. Trans-Canada Highway No. 1 passes through the middle of the Afton mining lease several kilometres west of its junction with the Coquihalla Highway No. 5. Access to the mine site is by a mine road located off of the Trans-Canada Highway.

New Gold’s holdings in the Project area comprise two groups, Afton and Ajax. The New Afton deposit lies within the Afton Group. The Afton Group consists of 44 mineral claims, a portion of which have been surveyed and are included in a mining lease issued on November 9, 2006. Total area of all claims and the lease is 7,352.7 ha.

New Gold holds surface rights on approximately 4,300 acres surrounding the New Afton Project, purchased from subsidiaries of Teck Resources Limited (Teck) on October 25, 2007. The terms of the purchase agreement provide title to fee simple land, a covenant to obtain a release of grazing leases, and access to surface improvements, including a water pipeline to the site from Kamloops Lake.

HISTORY

Exploration in the Afton area began in the mid-1800s, as prospectors pushed into the interior of British Columbia following the Fraser and Caribou gold rushes. The Iron Mask property, staked in 1896, was the first in the Kamloops district. The Afton property claims were staked over the Pothook workings in 1949. This was followed by sporadic, and largely unsuccessful, exploration work by a number of parties through the 1950s and 1960s.

In 1970, a private company called Afton Mines Ltd. (Afton Mines) obtained a drill intersection of 170 ft of 0.4% Cu from what ultimately became the Afton deposit. For the next three years, over 150,000 ft of drilling was carried out by a number of operators.

Teck Corporation and Iso Mines Ltd. bought acquired the Afton property in 1973, and initiated engineering and metallurgical studies. A production decision was taken in

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October 1975, with production commencing at the Afton open pit mine in late 1977.

At the start of production, the reserves were 34 million tons grading 1% Cu, 0.016 oz/ton Au, and 0.12 oz/ton Ag (30.8 million tonnes grading 1% Cu, 0.58 g/t Au, and 4.2 g/t Ag). Mining took place on the property at the Afton, Crescent, Pothook, and Ajax pits. The mine closed in 1991, reopened again in 1994, closing finally in 1997.

In 1999, the Afton mining leases expired and the ground was staked by two geologists. Later that year, DRC Resources Corporation (DRC) acquired an option on the property, surrounded it with additional staking, and carried out surface drilling in the New Afton deposit.

In February 2001, DRC completed a Scoping Study based on drilling results to date. Further definition drilling was recommended to confirm the continuity of the mineralization. An Advanced Scoping Study completed in February 2004 contemplated an underground mine, using a block caving method, feeding a conventional flotation mill operating at a rate of 9,000 tpd. Measured and Indicated Mineral Resources were 68.7 million tonnes grading 1.08% Cu, 0.85 g/t Au, and 2.62 g/t Ag.

In late 2004, a portal was collared on the south wall of the Afton pit, and 1,915 m of decline and drift were driven to provide access for definition drilling and bulk sampling of the deposit. Diamond drilling, both from surface and underground, has been carried out more or less continuously up to the time of writing of this report. At the cut-off date for the 2006 Mineral Resource estimate, the total amount of diamond drilling was 84,755 m in 182 holes.

DRC underwent a name change to New Gold Inc. in May 2005. New Gold commissioned a Feasibility Study from Hatch Ltd. in 2005. The study, completed in 2007, contemplated a block cave mine and conventional grinding/flotation mill operation with a daily throughput of 11,000 tonnes. Mineral Reserves comprised 44.4 million tonnes in the Probable category grading 0.98% Cu, 0.72 g/t Au, and 2.27 g/t Ag. Mine life was estimated to be 12 years, with preproduction capital costs of $268 million and Life of Mine (LOM) sustaining capital of $215 million

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Exploration drilling during 2006-2007 resulted in the discovery of the C Zone located below the main body of mineralization.

At the time of writing of this report, underground development was underway, the mill building had been erected with exterior cladding installed, and mill foundations had been poured. Surface construction was suspended until late 2010, and underground development was expected to be completed in 2012.

GEOLOGY AND MINERALIZATION

The Afton area is underlain primarily by rocks of the Upper to Lower Palaeozoic Quesnel Terrane, a package of island-arc sedimentary rocks, volcanic assemblages and associated intrusive phases. Quesnel Terrane was accreted onto the North American continental margin during the Early to Mid-Jurassic and is distributed throughout the entire length of British Columbia. It forms part of the Intermontane Belt, which is host to many porphyry Cu-Au-Mo deposits including Copper Mountain, Afton, Highland Valley, Mount Polley, Gibraltar, Kemess, and Galore Creek.

In the immediate Project area, the Quesnel Terrane comprises Late Triassic to Early Jurassic Nicola Group island-arc volcanic and sedimentary rocks and coeval alkalic to subalkalic intrusions of the Iron Mask batholith.

The New Afton deposit is hosted within the Cherry Creek member of the Iron Mask batholith complex. The Iron Mask complex is a multi-phase plutonic body exposed in a southeast-trending belt measuring 34 km long by 5 km wide. The Cherry Creek phase is a partially fault-bounded body trending east-northeast through the deposit area, curving to east side of the property to a more southeasterly trend. At Afton, the Cherry Creek intrusive is a variably and multiply brecciated assemblage of porphyritic and equigranular monzonite-monzodiorite. Cherry Creek rocks include fine- to medium-grained pyroxene-hornblende monzodiorite and fine-grained biotite-monzonite, varying to more dioritic composition, and minor syenite. The principal host phase of the Cherry Creek is a magnetic fine- to medium-grained porphyritic pyroxene-hornblende monzodiorite which forms a wedge of intrusive breccia between the Nicola and Pothook rocks. Contacts to the west and southwest are with Nicola Group volcanic rocks and to the east and southeast with the Pothook diorite.

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The Project area is traversed by regional-scale fault zones which are believed to be the principal control to intrusion of the batholithic rocks, as well as the mineralization processes. North- to northwest-trending steep to moderate-dipping fault zones are primary controlling structures for emplacement of the Iron Mask intrusives. East- to northeast-trending faults are important controls to hydrothermal alteration and mineralization. The faulting and associated fracturing also provided conduits for meteoric waters, which gave rise to weathering and supergene enrichment of the primary sulphide mineralization.

A steeply dipping body of serpentinized and sheared picrite has been entrained within the Hanging Wall Fault, a major structural control to the New Afton deposit. This unit separates strongly mineralized, pyrite-poor, potassically altered monzodiorite to the northwest from pyritic, sericite-carbonate-clay-chlorite-altered monzodiorite to the southeast. Ground conditions within the picrite are poor, which has hampered efforts to drill through it. As a result, drill information on the hanging wall side of the picrite is limited.

The bulk of the deposit occupies a large tabular mass measuring 900 m long by approximately 100 m wide and spanning a vertical distance of approximately 350 m. The zone dips vertically to steeply south-southeast and plunges at moderate angles to the southwest. Two subsidiary satellite bodies occur to the northeast of the Main Zone. One of these zones appears to branch off of the Main Zone and strikes in a somewhat more easterly trend, dipping steeply to the south-southeast. The other, which is further away from the Main Zone has a similar strike to the Main Zone but dips at a flatter angle (approximately 60° southeast).

The deposit is porphyry-style, with very distinct structural controls. The primary mineralizing event was preceded by development of disseminated magnetite-pyrite. Primary economic sulphide mineralization occurred in association with potassic alteration (principally K-spar) resulting in destruction of the earlier magnetite-pyrite assemblage. Following this, carbonate veinlets (principally ankerite) developed along and surrounding the faults, along with varying amounts of pyrite, clay gouge, and sericite.

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Subsequent faulting reshuffled portions of the deposit in a manner that is not completely understood. This has resulted in truncation of the zone along the Hanging Wall Fault in places, and apparent displacement of mineralized blocks. These blocks may occur as isolated pods in the hanging wall or may also be situated adjacent to the Main Zone such that they appear to comprise a continuous body of mineralization crossing the fault.

DRILLING

All drilling completed to date on the Project has been diamond core drilling, conducted in a series of programs over a period spanning 2000 to 2008. Most of the drilling has been for exploration and resource definition. However, in recent years (2007 and later), geotechnical holes have been drilled for groundwater monitoring and collection of data for mine design. To date, 277 holes have been drilled for a total of 130,439 m.

MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

Scott Wilson RPA carried out an audit of the most recent Mineral Resource estimate, which was prepared by Norwest Corporation (Norwest) in September 2009.

The Norwest Mineral Resource estimate is shown in Table 1-3.

TABLE 1-3 MINERAL RESOURCE SUMMARY – SEPTEMBER 2009

New Gold Inc. – New Afton Project

Tonnage (M tonnes)	Cu (%)	Au (g/t)	Ag (g/t)	Pd (g/t)
Measured
31.8	1.35	0.97	2.9	0.11
Indicated
20.0	1.10	0.82	2.8	0.11
Measured and Indicated
51.8	1.26	0.91	2.9	0.11
Inferred
12.8	0.98	0.70	2.4	0.09

Notes:

1.	CIM definitions were followed for Mineral Resources.

2.	Mineral Resources are estimated at a cut-off grade of 0.5% Cu.

3.	Mineral Resources are inclusive of Mineral Reserves.

As part of the audit, Scott Wilson RPA prepared an independent estimate for Cu, Au, Ag, and Pd and compared it to the Norwest estimate. A new probability shell was derived

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from an independently generated set of kriging parameters. This shell was then used to assign domains to the blocks and composites. Geostatistical analyses were carried out to derive variogram models, which were somewhat different from those generated by Norwest. Capping, search parameters, and classification criteria were kept the same as for the original models.

There is very little difference between the Norwest and Scott Wilson RPA estimates, which, in Scott Wilson RPA’s opinion, provides additional validation of the Norwest model. Scott Wilson RPA also notes that the global estimate appears to be quite insensitive to variations in either the probability shell or the variogram models.

Scott Wilson RPA recommends reporting the Mineral Resources at a CuEq cut-off of 0.4% CuEq. The Cu-equivalence calculation incorporates the value of the Au in the overall value of the resource body. In Scott Wilson RPA’s opinion, a 0.4% CuEq cut-off is supported by the present financial model for the mine. Table 1-4 summarizes the Mineral Resources estimate at the 0.4% CuEq cut-off.

TABLE 1-4 MINERAL RESOURCE SUMMARY – 0.4% CuEq CUT-OFF

New Gold Inc. – New Afton Project

Tonnage	Cu	Au	Ag	Pd
(M tonnes)	(%)	(g/t)	(g/t)	(g/t)
Measured
35.1	1.26	0.91	2.8	0.11
Indicated
29.9	0.85	0.67	2.2	0.10
Measured and Indicated
65.0	1.07	0.80	2.5	0.10
Inferred
25.2	0.66	0.54	1.7	0.08

Notes:

1.	CIM definitions were followed for Mineral Resources.

2.	Mineral Resources are estimated at a cut-off grade of 0.4% CuEq.

3.	Mineral Resources are estimated at US$2.50/lb Cu and US$900/oz Au.

4.	Mineral Resources are inclusive of Mineral Reserves.

5.	Cu-equivalence was determined using metal prices of US$2.50/lb Cu and US$900/oz Au, with no provision for metallurgical recovery.

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The Mineral Reserve estimate is shown in Table 1-5. The estimate is based on the resource block model prepared by New Gold and reviewed by Scott Wilson RPA.

TABLE 1-5 MINERAL RESERVE ESTIMATE – SEPTEMBER 2009

New Gold Inc. – New Afton Project

Category	Tonnage (M tonnes)	Cu (%)	Au (g/t)	Ag (g/t)
Probable	47.4	0.95	0.69	2.03

Notes:

1.	CIM definitions were followed for Mineral Reserves.

2.	Mineral Reserves are estimated at an NSR cut-off grade of $19/t.

3.	Mineral Reserves are estimated at US$2.00/lb Cu and US$800/oz Au and a C$/US$ exchange rate of 0.85.

4.	Metallurgical recoveries for copper vary from 69% to 89%.

MINING

The New Afton deposit is being developed as a block cave mining operation to produce 4 Mtpa of copper-gold ore for processing in a flotation plant for the recovery of copper in concentrates and gold in concentrates. Underground development is underway using New Gold crews which replaced the mining contractor crews and commenced development in February 2009. There is supporting infrastructure at the site including access, offices, staff, equipment, and a concrete batch plant. The new mill building has been erected and the grinding mills are mechanically installed. No significant surface construction was underway at the time of the site visit and the focus is on underground development and project planning and engineering. This report considers the development of the Project from the end of the second quarter of 2009 onwards.

Engineering studies have indicated that the New Afton deposit is amenable to block caving and the Mineral Reserves have been estimated through the application of Gemcom’s block cave modelling software (PCBC) to determine the mine production schedules. The deposit has been broken into three zones, B1, B2, and B3. The B1 and B2 blocks are closer to surface and will provide the initial mine production. Both the B1 and B2 blocks are expected to cave readily. The B3 block is deeper and the rock is more competent so that caving is marginal and it may be necessary to implement some additional work to assist the cave development.

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The B1 and B2 block undercut is approximately 120 m above the B3 undercut. At each block an undercut will be developed along with an extraction level. At the extraction level, drawpoints will be developed on a 13 m spacing. Ore will be mucked by load-haul-dump (LHD) units and hauled to ore passes to drop to a tramming level 14 m below, from where the ore will be trammed by large LHDs to the crusher. Ore from the B3 block will be hauled by 50-tonne truck to the crusher level. From the crusher the ore will be conveyed to the mill via a 4.5 km long conveyor system.

The mine is accessed by two portals for development, one in the open pit and the other being the conveyor portal near the mill. With the subsidence due to caving, the drift from the open pit will be lost and the secondary exit will be by conveyance in a ventilation borehole.

MINERAL PROCESSING

The New Afton mill has been designed to process 11,000 tpd, or 4 Mtpa, of ore at full capacity and recover copper, gold, and silver.

The mill will process a blend of primary, transition, and supergene ores. The process will utilize conventional crushing, grinding, and concentration processes to separate valuable minerals from gangue into concentrates at marketable grades. Mineral separation will be achieved by gravity concentration and differential flotation. A regrinding stage will be included in the flotation circuit.

The tailings handling system will pump the combined tailings (from the rougher flotation circuit and the cleaner scavenger circuit) to the Cyclone Plant, a set of primary cyclones located adjacent to the tailings storage facility (TSF).

The cyclone underflow product will be delivered to a skid-mounted secondary cyclone pack located on the top of the tailings dam. The sand or coarse fraction from these secondary cyclones will be used to construct the tailings dam and the secondary cyclone overflow product will be used to seal the wet side of that dam. A centreline construction method will be used to establish the tailings dam structure. There will be three seepage collection ponds below the dry side of the main tailings dam to recover and return seepage back to the TSF.

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ENVIRONMENTAL

Much of the Project area has been greatly altered such that natural undulating has been replaced by level rock dumps, open pits, and severely compacted borrow areas. The dominant water bodies are now related to mining-related impoundments. The baseline environmental studies undertaken considered both the natural and anthropogenic landscapes. Environmental baseline studies were undertaken on the Project site between April and December 2006. All baseline studies were implemented in accordance with provincial data collection guidelines where applicable, and generally accepted scientific procedures.

On October 31, 2007, the Ministry of Energy, Mines and Petroleum Resources (MEMPR) issued Mine Permit M229 approving the work system and reclamation program in respect of the New Afton Project. The Mine Permit obligates the New Gold to post reclamation security in the aggregate amount of $9.5 million. The initial installment of $3.5 million was posted by the New Gold.

New Gold has entered into an agreement with local First Nations to pay a royalty on the value of the product with a set minimum annual royalty.

CAPITAL AND OPERATING COSTS

The remaining preproduction capital budget for the Project is C$460.9 million in Q2 2009 Canadian dollars. Sunk costs are not included in the capital summary. Capital cost estimates have been prepared by AMEC, New Gold, and AMC Mining Consultants (Canada) (AMC) in the course of the detailed engineering work which was being done by AMEC until mid-2009.

Capital cost estimate revisions and adjustments continue as Project optimization tasks are still underway. The mine capital costs are built up from first principles for the mine development activities together with quotations for major equipment items. The sustaining capital estimate for the life of the mine is C$172.3 million.

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The life of mine operating cost is estimated to be C$17.27 per tonne milled. The costs do not include concentrate shipping or treatment, nor do they include corporate overheads. Operating costs have been compiled by New Gold and AMC. The costs for all areas were built up from first principles with a high level of detail in the cost build-up.

PROJECT EXECUTION PLAN

Development of the New Afton Project is underway, with the underground development being undertaken by company crews. A Project team is in place with administrative, technical (geology, mine and mill) and mine development personnel on staff. There are also engineering consultants supporting the group.

The current Project schedule is not in a finished form; this is an aspect of the Project planning which Scott Wilson RPA recommends be addressed to generate a comprehensive Project development schedule including all of the changes in the Project since the 2007 Feasibility Study.

Mine development is scheduled to continue until and after the commencement of production. The development of the B3 zone has been deferred to reduce the upfront capital costs. Mill construction is forecast to recommence at the start of 2011, with the mill planned to be in commissioning mode in Q2 of 2012 and operating at the start of Q3 2012.

New Gold will continue to use AMEC for engineering and procurement, but will do construction management for the remaining mill construction packages with its own team, led by a recently hired construction manager.

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2 INTRODUCTION

Scott Wilson Roscoe Postle Associates Inc. (Scott Wilson RPA) was retained by Mr. Mark Petersen, Vice President, Exploration, of New Gold Inc. (New Gold), to prepare an independent Technical Report and Feasibility Update on the New Afton Cu-Au-Ag-Pd Project (the Project) near Kamloops, British Columbia, Canada. The purpose of this report is to support the public disclosure of updated Mineral Resources and Mineral Reserves for the Project. This Technical Report conforms to NI 43-101 Standards of Disclosure for Mineral Projects (NI 43-101). Scott Wilson RPA visited the property on September 3 and 4, 2009.

New Gold is a publicly traded mining company, focused on gold production, with operations in the United States, Mexico, and Australia. New Gold’s corporate offices are located in Vancouver, British Columbia and Toronto, Ontario. The company is listed on Toronto (TSX:NGD) and New York (NYSE: AMEX) stock exchanges.

New Afton is a former producer which operated an open pit mine from 1977 to 1987. New Gold acquired an option on the New Afton property in 1999 and began exploration work in 2000. The Project is in the development stage and is a planned 11,000 tpd (nominal) block cave operation scheduled to commence milling in mid-2012. Copper and gold concentrates will be produced at the Project. Development commenced in 2007 but, except for engineering and other limited activities, was halted due to the global economic conditions in the third quarter of 2008. Underground development was underway at the time of the site visit and surface construction is scheduled to restart in 2010.

Currently, the major assets and facilities associated with the Project are:

·	The resources and reserves of the New Afton copper gold deposit.

·	Site access from the Trans-Canada Highway which is immediately adjacent to the site.

·	Underground mine workings including a conveyor gallery, a second access from the open pit, and ventilation raises to surface.

·	Facilities providing basic infrastructure to the mine, including electric power, heat, water supply, communications, and mine offices.

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·	An underground mobile equipment fleet and support facilities at both the upper, conveyor gallery portal and the portal in the open pit.

·	Dewatering facilities for pumping from the open pit to a storage area in Pothook Lake, which has been bermed to increase its storage capacity.

·	Construction support equipment including a concrete batch plant and trucks for the delivery of concrete and surface support equipment.

·	Engineering and administration offices, mine dry and first aid facilities.

·	A new concentrator building with the semi-autogenous grinding (SAG) mill, ball mill, and vertical mill mechanically installed.

·	Various mill equipment parts stored on site in preparation for installation.

·	A partially cleared and prepared tailings storage facility.

·	Site access roads.

·	A mine development permit issued by the British Columbia government.

Prior Scott Wilson RPA involvement is as follows:

·	Conducted an audit of the Mineral Resource estimate; issued a Technical Report in 2006 (Wallis and Giroux, 2006).

·	Prepared a Mineral Resource estimate for the Project in 2006 and updated the estimate in 2007.

·	Participated in a Feasibility Study for the Project (Mineral Resources); Technical Report issued April 2007 (Hatch, 2007).

SOURCES OF INFORMATION

Site visits were carried out by David Rennie, P.Eng., Scott Wilson RPA Principal Geologist, and Dennis Bergen, P.Eng., Scott Wilson RPA Associate Principal Mining Engineer.

Discussions were held with personnel from New Gold:

·	Mr. David Nicholls, Manager of Technical Services

·	Mr. Nick Kwong, Project Engineer

·	Mr. Eric Strom, Senior Mine Engineer

·	Mr. Craig Lockhart, Maintenance and Surface Manager

·	Ms. Jane McCaw, Database and Land Specialist

·	Mr. Daniel Kahler, AMC Consultants

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·	Mr. Richard Hall, Geologist

·	Mr. Brian Bower, Geologist

The geology and Mineral Resource sections of this report were prepared by David Rennie, P. Eng. Metallurgy sections were prepared by Kevin Scott, P. Eng., Principal Metallurgist for Scott Wilson RPA. The mining, permitting, conversion of mineral resources to reserves, and financial model sections were prepared by Dennis Bergen, P. Eng., Associate Principal Mining Engineer for Scott Wilson RPA.

The documentation reviewed, and other sources of information, are listed at the end of this report in Section 22, References.

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SCOTT WILSON RPA WWW.SCOTTWILSON.COM

LIST OF ABBREVIATIONS

Units of measurement used in this report conform to the SI (metric) system. All currency in this report is Canadian dollars (C$) unless otherwise noted.

m	micron	kPa	kilopascal
°C	degree Celsius	kVA	kilovolt-amperes
°F	degree Fahrenheit	kW	kilowatt
mg	microgram	kWh	kilowatt-hour
A	ampere	L	litre
a	annum	L/s	litres per second
bbl	barrels	m	metre
Btu	British thermal units	M	mega (million)
C$	Canadian dollars	m2	square metre
cal	calorie	m3	cubic metre
cfm	cubic feet per minute	min	minute
cm	centimetre	MASL	metres above sea level
cm2	square centimetre	mm	millimetre
d	day	mph	miles per hour
dia.	diameter	MVA	megavolt-amperes
dmt	dry metric tonne	MW	megawatt
dwt	dead-weight ton	MWh	megawatt-hour
ft	foot	m3/h	cubic metres per hour
ft/s	foot per second	opt, oz/st	ounce per short ton
ft2	square foot	oz	Troy ounce (31.1035g)
ft3	cubic foot	oz/dmt	ounce per dry metric tonne
g	gram	ppm	part per million
G	giga (billion)	psia	pound per square inch absolute
Gal	Imperial gallon	psig	pound per square inch gauge
g/L	gram per litre	RL	relative elevation
g/t	gram per tonne	s	second
gpm	Imperial gallons per minute	st	short ton
gr/ft3	grain per cubic foot	stpa	short ton per year
gr/m3	grain per cubic metre	stpd	short ton per day
hr	hour	t	metric tonne
ha	hectare	tpa	metric tonne per year
hp	horsepower	tpd	metric tonne per day
in	inch	US$	United States dollar
in2	square inch	USg	United States gallon
J	joule	USgpm	US gallon per minute
k	kilo (thousand)	V	volt
kcal	kilocalorie	W	watt
kg	kilogram	wmt	wet metric tonne
km	kilometre	yd3	cubic yard
km/h	kilometre per hour	yr	year
km2	square kilometre

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3 RELIANCE ON OTHER EXPERTS

This report has been prepared by Scott Wilson Roscoe Postle Associates Inc. (Scott Wilson RPA) for New Gold Inc. (New Gold). The information, conclusions, opinions, and estimates contained herein are based on:

·	Information available to Scott Wilson RPA at the time of preparation of this report,

·	Assumptions, conditions, and qualifications as set forth in this report, and

·	Data, reports, and other information supplied by New Gold and other third party sources.

For the purpose of this report, Scott Wilson RPA has relied on ownership information provided by New Gold. Scott Wilson RPA has not researched property title or mineral rights for the New Afton Project and expresses no opinion as to the ownership status of the property.

Scott Wilson RPA has relied on New Gold for guidance on applicable taxes, royalties, and other government levies or interests, applicable to revenue or income from the New Afton Project

Except for the purposes legislated under provincial securities laws, any use of this report by any third party is at that party’s sole risk.

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SCOTT WILSON RPA WWW.SCOTTWILSON.COM

4 PROPERTY DESCRIPTION AND LOCATION

The New Afton Project is located at Latitude 50°39’ north and Longitude 120°31’ west, approximately 350 km northeast of Vancouver and 10 km west of the City of Kamloops, in the South-Central Interior of British Columbia (Figure 4-1). New Gold’s holdings in the Project area comprise two groups, Afton and Ajax. The New Afton deposit lies within the Afton Group. The Afton Group consists of 44 mineral claims that have been converted to the new grid staking system (Figure 4-2). A list of these claims is provided in Table 4-1. A portion of the Afton claims have been surveyed and are included in a mining lease, which was issued on November 9, 2006 (Figure 4-2). Total area of all claims and the lease is 7,352.7 ha.

The project survey control is based on the Universal Transverse Mercator (UTM) coordinate system. It is based on the Zone 10 North projection, using the World Geodetic System 1984 (WGS’84) datum. The UTM coordinates place the New Afton project at 5,614,800N and 675,500E at a surface elevation of 700 MASL.

LAND TENURE

New Gold holds surface rights for approximately 4,300 acres surrounding the New Afton Project, purchased from subsidiaries of Teck Resources Limited (Teck) on October 25, 2007. The terms of the purchase agreement provide title to fee simple land, a covenant to obtain a release of grazing leases, and access to surface improvements, including a water pipeline to the site from Kamloops Lake.

Reclamation permits covering the existing pits, plant facility and tailings pond, issued by the British Columbia Ministry of Energy and Mines, are held in the name of the previous operator, Teck.

Under an Option to Purchase Agreement (“the Option”) dated September 22, 1999, DRC Resources Corporation (DRC), the predecessor company to New Gold, acquired the exclusive right for 90 days to purchase a 100% undivided working interest in the Afton 1 – 11 (the “Original Claims”) as to 50% from Westridge Enterprises Ltd., a non-reporting

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British Columbia company wholly owned by John H. Kruzick, a director of DRC, and as to 50% from Indo-Gold Development Ltd., a non-reporting British Columbia company owned by John Ball, a geologist (the “Optionors”). The Option provided for consideration to be a 10% Net Profit Royalty (NPR) and a property management agreement with the Optionors. The terms of the agreement required a series of payments in shares and expenditure, totalling 2 million shares and $6.5 million in expenditures over a nine year time frame. The payments and work expenditures have been completed as required and New Gold now owns the property subject to a 10% NPR.

On June 27, 2002, New Gold acquired an option exercisable on or before December 1, 2010, to purchase the Optionors’ 10% NPR for a cash consideration of $2 million and the Option was amended by removing the exclusive right of the Optionors to act as the operators of the New Afton Cu-Au Project. In addition to the Original Claims, New Gold has staked other claims for its own account to consolidate its land position at the Project.

TABLE 4-1 MINERAL TENURES

New Gold Inc. - New Afton Project

Tenure No.	Claim Name	Map No	Good To	Area (ha)	Tag No	Status	MTO Area
372644	AFTON 8	092I068	2017/Mar/08	9.7	686853M	Legacy	25.000
372645	AFTON 9	092I068	2017/Mar/08	1.0	686854M	Legacy	25.000
372646	AFTON 10	092I068	2017/Mar/08	9.0	686855M	Legacy	25.000
372647	AFTON 11	092I068	2017/Mar/08	4.9	686856M	Legacy	25.000
378688	AFTON 8	092I068	2017/Mar/08	432.4	220718	Legacy	500.000
379304	AFTON 19	092I068	2017/Mar/08	0.6	694839M	Legacy	25.000
379305	AFTON 20	092I068	2017/Mar/08	0.0	694838M	Legacy	25.000
379306	AFTON 25	092I068	2017/Mar/08	0.0	700605M	Legacy	25.000
379311	AFTON 30	092I068	2017/Mar/08	0.0	700603M	Legacy	25.000
379312	AFTON 31	092I068	2017/Mar/08	0.0	694806M	Legacy	25.000
514167	AFTON	092I	2017/Mar/08	225.1		Cell	225.139
514194		092I	2017/Mar/08	1,520.8		Cell	1,637.761
517047	AFTON	092I	2017/Mar/08	21.1		Cell	40.974
517157	AFTON	092I	2017/Mar/08	200.8		Cell	204.764
517360	NEW AFTON	092I	2017/Mar/08	12.4		Cell	20.492
524303	AFTON DAM	092I	2017/Mar/08	0.0		Cell	20.492
524304	AFTON DAM 1	092I	2017/Mar/08	7.4		Cell	20.491
524305	AFTON DAM 2	092I	2017/Mar/08	5.2		Cell	20.491
525508	AFTON DAM 3	092I	2017/Mar/08	1.0		Cell	20.492

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Tenure No.	Claim Name	Map No	Good To	Area (ha)	Tag No	Status	MTO Area
528243	SMELTER	092I	2017/Mar/08	15.0		Cell	20.485
528575	KAMLOOPS COPPER PAST	092I	2011/Mar/08	4.5		Cell	81.772
529020	COPPER LOAD	092I	2010/Feb/27	17.1		Cell	20.485
529257	MAXINE NO 2	092I	2011/Mar/08	3.6		Cell	40.881
529749	MAXINE GROUP #2	092I	2011/Mar/08	6.0		Cell	61.326
529830	MAXINE #3	092I	2011/Mar/08	0.2		Cell	102.218
534320	KAMLOOPS GOLD DEPOSIT	092I	2011/Mar/08	118.447		Cell	122.762
534321	KAMLOOPS GOLD NO 2	092I	2011/Mar/08	35.943		Cell	40.926
534322	JUST IN CASE	092I	2011/Mar/08	18.738		Cell	20.467
534787	AF EXT 11	092I	2017/Mar/08	4.5		Cell	20.482
534788	AF EXT12	092I	2017/Mar/08	17.1		Cell	20.476
534795	AF EXT 13	092I	2017/Mar/08	3.6		Cell	20.482
534896	AFTON N 2	092I	2017/Mar/08	204.853		Cell	224.906
534923	MAX\1	092I	2011/Mar/08	122.681		Cell	122.681
536450	KAMLOOPS BIG ONE	092I	2011/Mar/08	122.794		Cell	122.794
537230	AFTON DAM 3	092I	2017/Mar/08	6.0		Cell	40.983
537231	AFTON DAM 2	092I	2017/Mar/08	0.2		Cell	40.978
537385	AFTON NW 2	092I	2017/Mar/08	511.249		Cell	511.249
537387	AFTON NW 3	092I	2017/Mar/08	981.858		Cell	981.858
538320	AFTON NW 1	092I	2017/Mar/08	901.549		Cell	1,105.178
548542	AFTON NW 4	092I	2009/Mar/08	40.9191		Cell	40.9191
549226	AFTON NW 5	092I	2009/Mar/08	453.138		Cell	491.2601
549268	AFTON NW 6	092I	2009/Mar/08	13.094		Cell	20.4746
549270	AFTON NW 7	092I	2009/Mar/08	33.098		Cell	81.9232
552399	ML EXT 1	092I	2010/Feb/20	1.132		Cell	20.4819
552400	ML EXT 2	092I	2010/Feb/20	5.197		Cell	20.4834
554986	AFTON NW 8	092I	2009/Mar/08	163.5743		Cell	163.5743
594462	AFTON EEA	092I	2009/Nov/18	184.41		Cell	184.41
546063		092I	2010/Nov/29	902.3		ML	902.300
606247	AFTON NW 8	092I	2010/Jun/18	8.656		Cell	81.940

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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

ACCESSIBILITY

The New Afton Project is located 350 km northeast of Vancouver and 10 km west of Kamloops, British Columbia, in the South-Central Interior of British Columbia, on the site of the past producing Afton Mine, a previous operation of Teck. The Project is located on the south side of the Thompson River Valley. Trans-Canada Highway No. 1 passes, several kilometres west of its junction with the Coquihalla Highway No. 5, through the middle of the Afton mining lease. Access to the mine site is by a mine road located off of the Trans-Canada Highway.

CLIMATE

The Kamloops area is in the rain shadow of the Coast Mountains. Precipitation is minor, averaging about 257 mm annually (of which 175 mm is rainfall), with light winter snow and infrequent rain in the spring and fall. The area has warm summers where temperatures can reach 38°C and cool winters where temperatures hover around the freezing mark. During the winter, short periods of cold weather can occur where temperatures drop to as low as -29°C.

LOCAL RESOURCES

Kamloops is a major transportation hub for highway, air, and railroad. Forestry, ranching, mining, and tourism are the most important economic activities in the area. The city has an airport with daily air service from Vancouver, British Columbia, and Calgary, Alberta, and is serviced by both Canadian National and Canadian Pacific Railways. Kamloops is a natural resource-based city of 80,000 people. The area has a ready supply of trained workers and professionals and has suppliers and contractors to support heavy industry.

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INFRASTRUCTURE

BC Hydro transmission lines, a Terasen natural gas pipeline, and a Pembina oil pipeline traverse the mining lease north of the Afton pit. A water pipeline approximately four kilometres in length can deliver fresh water from Kamloops Lake to the mine site. New Gold purchased the pipeline and pump house facilities from Teck as part of the purchase agreement in 2007. New Gold has obtained a temporary water permit and applied for a permanent water permit to withdraw water from Kamloops Lake for mining and milling operations.

The surface and underground infrastructure at the New Afton Project includes:

·	Site access from the Trans-Canada Highway which is immediately adjacent to the site.

·	Underground mine workings a conveyor gallery, a second access from the open pit, and ventilation raises to surface.

·	Facilities providing basic infrastructure to the site including electric power, water supply, communications, and mine offices.

·	An underground mobile equipment fleet and support facilities at both the upper, conveyor gallery portal and the portal in the open pit.

·	Dewatering facilities for pumping from the open pit to a storage area in Pothook Lake, which has been bermed to increase its volume.

·	Construction support equipment including a concrete batch plant and trucks for the delivery of concrete and surface support equipment.

·	Engineering and administration offices, mine dry and first aid facilities.

·	A new concentrator building with the SAG mill, ball mill and vertical mill mechanically installed.

·	A partially cleared and prepared tailings storage facility.

The old Teck concentrator building remains in its location on the east side of the open pit, but the New Afton facilities are located to the west of the pit. Similarly, the old Teck tailings storage area remains on site but is not planned for use by New Gold. Both the concentrator building and tailings pond are on land owned by others.

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PHYSIOGRAPHY

The landscape is characterized by hilly, till-covered, drumlinoidal terrain and dispersed, small, alkaline water bodies. Relief adjacent to Kamloops Lake is a few hundred metres or more. The most significant topographic features within the mining lease are the Afton and Pothook open pits and the reclaimed waste rock dumps of the Afton Operating Corporation, the former operator. Kamloops Lake is located north of the mining lease and bisects the Afton mineral tenure. Vegetation consists of open grasslands and sparse pine forests as a consequence of a continental, semi-arid climate.

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6 HISTORY

Exploration in the Afton area began in the mid-1800s, as prospectors pushed into the interior of British Columbia following the Fraser and Caribou gold rushes. The Iron Mask property, staked in 1896, was the first in the Kamloops district. A 100 ft shaft was sunk on the Pothook deposit in 1898. Mining was carried out from the turn of the century through until 1927 at several gold, copper, and silver mines including the Pothook, Iron King, Copper King, and Iron Mask. The Afton property claims were staked over the Pothook workings in 1949 by Axel Bergland. This was followed by sporadic, and largely unsuccessful, exploration work by a number of parties through the 1950s and 1960s.

Mr. Chester Millar acquired the property in the mid-1960s and formed a private company called Afton Mines Ltd. (Afton Mines) to carry out exploration work. In 1970, Afton Mines obtained a drill intersection of 170 ft of 0.4% Cu from what ultimately became the Afton deposit. For the next three years, over 150,000 ft of drilling was carried out by a number of operators. Duval Corporation and Quintana Minerals took options on the property in 1970 but dropped them in 1971. They were followed in 1972 by Canex Placer. Also in 1972, Teck Corporation and Iso Mines Ltd. (Iso) purchased an equity interest in Afton Mines.

Teck and Iso bought Canex Placer’s interest for $4.0 million in 1973, and initiated engineering and metallurgical studies. A production decision was taken in October 1975, with production commencing at the Afton open pit mine in late 1977. At the start of production, the reserves were 34 million tons grading 1% Cu, 0.016 oz/ton Au, and 0.12 oz/ton Ag (30.8 million tonnes grading 1% Cu, 0.58 g/t Au, and 4.2 g/t Ag). Mining took place on the property at the Afton, Crescent, Pothook, and Ajax pits. The mine closed in 1991, reopened again in 1994, closing finally in 1997.

Five deep diamond drill holes, drilled in 1980 below and to the southwest of the Afton pit, pierced what is now referred to as the New Afton deposit. Teck carried out a study to determine the feasibility of mining this zone from underground, but shelved the Project.

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In 1999, the Afton mining leases expired and the ground was staked by Westridge Ltd., and Indogold Development Ltd. DRC Resources Corporation (DRC) acquired an option on the property and surrounded it with additional staking. The following year, DRC began exploration work with 9,320 m of surface diamond drilling in 21 NQ holes in the New Afton deposit.

In February 2001, DRC completed a Scoping Study based on drilling results to date. The Scoping Study indicated that the New Afton deposit could be profitably exploited, and this led to further definition drilling to confirm the continuity of the mineralization. The successful conclusion of this phase of exploration led to the commissioning of an Advanced Scoping Study in 2003 and completed in February 2004 (Behre Dolbear, 2004). The study contemplated an underground mine, using a block caving method, feeding a conventional flotation mill operating at a rate of 9,000 tpd. Measured and Indicated Mineral Resources were 68.7 million tonnes grading 1.08% Cu, 0.85 g/t Au, and 2.62 g/t Ag. The report estimated an after-tax internal rate of return (IRR) for the Project of 19.9%, with a payback of capital in 3.7 years.

In late 2004, a portal was collared on the south wall of the Afton pit, and 1,915 m of decline and drift were driven to provide access for definition drilling and bulk sampling of the deposit. The drifting was completed by September 2005. Diamond drilling, both from surface and underground, has been carried out more or less continuously up to the time of writing of this report. At the cut-off date for the 2006 Mineral Resource estimate, the total amount of diamond drilling was 84,755 m in 182 holes.

DRC underwent a name change to New Gold Inc. in May 2005. Total exploration expenditures on the New Afton Project by DRC and New Gold to the end of the 3rd quarter of 2006 were C$37.2 million.

New Gold commissioned a Feasibility Study for the Project at the end of 2005, and the results of this study were released in April 2007 (Hatch, 2007). The study contemplated a block cave mine and conventional grinding/flotation mill operation with a daily throughput of 11,000 tonnes. Mineral Reserves comprised 44.4 million tonnes in the Probable category grading 0.98% Cu, 0.72 g/t Au, and 2.27 g/t Ag. Mine life was

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estimated to be 12 years, with preproduction capital costs of $268 million and Life of Mine (LOM) sustaining capital of $215 million. Pre-tax IRR, using base-case metal prices of US$2.01/lb Cu, US$487/oz Au, and US$8.54/oz Ag, was estimated to be 13.6%.

Also during the period 2006-07, exploration drilling resulted in the discovery of the C Zone located below the main body of mineralization.

At the time of writing of this report, underground development was underway, the mill building had been erected with exterior cladding installed, mill foundations had been poured, and the SAG Mill, Ball Mill and Verti-Mill installed. Surface construction was suspended until late 2010, and underground development was expected to be completed in 2012. This concept will allow the mill to be completed in time for the underground to deliver ore at the mill’s rated capacity of 11,000 tpd.

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7 GEOLOGICAL SETTING

REGIONAL GEOLOGY

The Afton area is underlain primarily by rocks of the Upper to Lower Palaeozoic Quesnel Terrane, a package of island-arc sedimentary rocks, volcanic assemblages and associated intrusive phases. Quesnel Terrane was accreted onto the North American continental margin during the Early to Mid-Jurassic, and is distributed throughout the entire length of British Columbia. It forms part of the Intermontane Belt, which is host to many porphyry Cu-Au-Mo deposits including Copper Mountain, Afton, Highland Valley, Mount Polley, Gibraltar, Kemess, and Galore Creek.

In the immediate Project area, the Quesnel Terrane comprises Late Triassic to Early Jurassic Nicola Group island-arc volcanic and sedimentary rocks, and coeval alkalic to subalkalic intrusions of the Iron Mask batholith (Figure 7-1). The Nicola Group consists of submarine and subaerial volcanic and sedimentary rocks. Volcanic components are intermediate to basic flows, with associated breccias and volcaniclastic units. The Nicola Group rocks have been regionally metamorphosed to lower greenschist facies.

Intruding the Nicola Group are Iron Mask diorite, monzodiorite, and monzonite plutons occurring along a northwest trend. Four principal phases of the Iron Mask batholith have been identified and these are the Pothook phase agmatite intrusion breccia, Pothook diorite, Cherry Creek Suite monzodiorite to monzonite, and Sugarloaf diorite. Near the margins of the Iron Mask batholith, Nicola Group rocks have been thermally metamorphosed to hornfels. The Iron Mask intrusive rocks are the principal hosts for Cu-Au-Ag mineralization at New Afton.

Post-mineralization Middle Eocene volcanic and sedimentary rocks of the Kamloops Group unconformably overlie the Nicola and Iron Mask units. These rocks were partially responsible for shielding the New Afton deposit from removal by erosion and glaciation.

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primary controlling structures for emplacement of the Iron Mask intrusives. East- to northeast-trending faults are important controls to hydrothermal alteration and mineralization. The faulting and associated fracturing also provided conduits for meteoric waters, which gave rise to weathering and supergene enrichment of the primary sulphide mineralization.

PROPERTY GEOLOGY

The New Afton deposit is hosted within the Cherry Creek member of the Iron Mask batholith complex (Figure 7-2). The Iron Mask complex is a multi-phase plutonic body exposed in a southeast-trending belt measuring 34 km long by 5 km wide. The Cherry Creek phase is the principal host unit for the New Afton deposit. It is a partially fault-bounded body trending east-northeast through the deposit area, curving to the east side of the property to a more southeasterly trend (Figure 7-3). At Afton, the Cherry Creek intrusive is a variably and multiply brecciated assemblage of porphyritic and equigranular monzonite-monzodiorite. Cherry Creek rocks include fine- to medium-grained pyroxene-hornblende monzodiorite and fine-grained biotite-monzonite, varying to more dioritic composition, and minor syenite. The principal host phase of the Cherry Creek is a magnetic fine- to medium-grained porphyritic pyroxene-hornblende monzodiorite which forms a wedge of intrusive breccia between the Nicola and Pothook rocks. Contacts to the west and southwest are with Nicola Group volcanic rocks and to the east and southeast with the Pothook diorite.

Contact relationships between various intrusive phases are complex and not completely understood. Cherry Creek rocks are difficult to discriminate from Pothook biotite-pyroxene diorite owing to overprinting by alteration. Strong potassic alteration occurs along the contact between the Pothook and Cherry Creek phases. Pothook rocks outcrop to the east and south of the Afton deposit. They are weakly porphyritic to equigranular and, where not altered, moderately to strongly magnetic. Poikiolitic biotite is diagnostic of the Pothook diorite.

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Nicola Group rocks border the Cherry Creek intrusive on west and southwest of the deposit but are not hosts to the economic mineralization. They consist of augite porphyry, polylithic volcanic breccias, and picrite flows and breccias. Xenoliths of Nicola rocks occur within the Cherry Creek intrusive, and on the southwest end of the deposit the Nicola form a large roof pendant overlying the Cherry Creek. A possible volcanic vent breccia phase appears to crosscut the Nicola strata

A steeply dipping body of serpentinized and sheared picrite has been entrained within the Hanging Wall Fault, a major structural control to the New Afton deposit. This unit separates separating strongly mineralized, pyrite-poor, potassically altered monzodiorite to the northwest from pyritic, sericite-carbonate-clay-chlorite-altered monzodiorite to the southeast. Ground conditions within the picrite are poor, which has hampered efforts to drill through it. As a result, drill information on the hanging wall side of the picrite is limited.

New Gold geologists have recently discerned that a late, post-mineral intrusive phase occupies a corridor oriented roughly parallel to and lying within the orebody. This intrusive post-dates the Cu-Au mineralization and is largely barren, except for scattered occurrences of limited lateral extent which may, in fact, be entrained blocks of earlier mineralized intrusive material. It has been called the “De-Mineralized Zone” (DMZ) by mine staff.

The property is traversed by mafic pyroxene-feldspar porphyry, monzonite, hornblende diorite porphyry, and hornblende-feldspar-quartz porphyry (latite) dykes. The dykes are late phases of the intrusive and difficult to discriminate from Cherry Creek rocks in places. They are post-mineral events, although some have pyrite-chalcopyrite mineralization along their margins.

Mesozoic units (i.e., Nicola and Cherry Creek) are unconformably overlain by Eocene Kamloops Group rocks. In the Project area, the Kamloops Group comprises coarse- to fine-grained clastic sedimentary rocks, along with a mafic flow. Coal beds within this package are visible in the north wall of the Afton Pit.

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Structure is a dominant feature of the geology of the deposit and has influenced all aspects of the host rock lithology, mineralizing events, and post-mineral processes. Zones of intersection between deep-seated northwest-southeast-trending faults and east-northeast fractures were primary controls to localization of the Triassic intrusive bodies (i.e., the Iron Mask batholith) and Nicola volcanic centres. The northwest-southeast structures are major regional fault zones believed to date back to Triassic times with repeated stages of reactivation through to the Tertiary.

East-northeast-striking steeply dipping fault zones were the primary control to the New Afton mineralization. Two faults, termed the Hanging Wall (east) and Footwall (west) faults, constrain the New Afton deposit to a relatively narrow steeply dipping corridor. They, along with moderately to steeply dipping east-west, southwest-northeast, and north-south fault zones have been mapped in the Afton pit (see Figure 7-2), and were subsequently picked up in diamond drill holes and underground mapping. These structures are thought to have controlled the flow of mineralizing hydrothermal fluids responsible for the alteration and deposition of sulphides. The pattern of intersection of the fault planes has imparted a southwest plunge to the New Afton zone.

Subsequent fault movements shuffled portions of the deposit, in parts truncating it, and transposing slices of waste and ore in others. The faults and subsidiary fracturing then provided conduits for weathering solutions, which resulted in supergene alteration down to relatively deep elevations (approximately 400 m below surface in places).

Lastly, during the Tertiary, tensional stress that developed in a northwest-southeast direction resulted in northwest-southeast-trending graben structures and the development of depositional basins for Kamloops Group sediments.

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8 DEPOSIT TYPES

The Afton deposits are Cu-Au silica-saturated, alkalic porphyry style deposits. Mineralization results from late stage hydrothermal activity driven by remnant heat within an alkalic intrusive complex. Thermal gradients within these systems give rise to broadly concentric, although often complexly intermingled, zones of alteration and mineralization. The distribution of alteration and mineral facies are largely influenced by dikes, veins, and fracture systems which concentrate and control fluid flow.

The New Afton deposit is unique in the sense that it is very strongly constrained by structure. The mineralization occurs as a tabular body that occupies a largely fault-bounded corridor, which traverses the property in an east-northeast/west-southwest direction (see Figure 7-2). This corridor is bounded on either side by the Hanging Wall (SE) and Footwall (NW) faults. The orebody does not come in contact with the Footwall Fault, however, mineralization is partially entrained within and likely disrupted by portions of the Hanging Wall Fault. Pods of mineralization are also known to occur outboard of the Hanging Wall Fault.

Very late epithermal processes have overprinted the porphyry system and introduced mercury, antimony, and arsenic in geochemically significant concentrations.

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9 MINERALIZATION

Hypogene mineralization comprises fine-grained disseminated chalcopyrite and relatively minor bornite or chalcocite. In weathered portions of the deposit, the chalcopyrite has been altered to bornite and chalcocite, and in the most intensely oxidized areas, native copper. Supergene mineralization is broadly constrained to the upper portions of the deposit, reaching to deeper areas along fracture zones.

Native gold is rarely observed. Gold and silver reportedly occur as electrum grains within the chalcopyrite and bornite.

Significant amounts of mercury, antimony, and arsenic occur in and around the deposit but are not thought to be directly related to the porphyry event(s). Rather, they are more likely to have been introduced during much later epithermal activity, similar to several other localities in the district.

The central core of mineralization comprises an early potassic alteration phase which includes intense biotite ± silica ± magnetite hornfels of Nicola Group and Cherry Creek monzonite. The earliest stages of the mineralization also include stockworks of magnetite + apatite + actinolite + biotite + calcite + chalcopyrite ± pyrite veins. Overprinted on the hornfels and magnetite facies is fracture-controlled to pervasive hydrothermal potassic alteration (K-feldspar + biotite + actinolite) and sulphide mineralization consisting of disseminated and veinlet-hosted chalcopyrite and pyrite with subsidiary bornite. Copper-gold mineralization occurs in association with strong potassic alteration and in crackle breccias proximal to magnetite, pyrite, and quartz-carbonate stockworks.

Pyrite stockworks encompass the potassic zones in a shell of phyllic alteration, which includes dolomite + sericite in quartz-carbonate stockworks associated with high-angle faults. This imparts a bleached appearance to the rock in contrast with the darker brownish-coloured hornfels.

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The latest stage of hydrothermal alteration associated with the primary mineralization event is a propylitic overprint consisting of actinolite + chlorite + epidote ± carbonate ± pyrite ± chalcopyrite.

The bulk of the deposit occupies a large tabular mass measuring 900 m long by approximately 100 m wide and spanning a vertical distance of about 350 m. The zone dips vertically to steeply south-southeast, and plunges at moderate angles to the southwest. Two subsidiary satellite bodies occur to the northeast of the Main Zone. One of these zones appears to branch off of the Main Zone and strikes in a somewhat more easterly trend, dipping steeply to the south-southeast. The other, which is further away from the Main Zone has a similar strike to the Main Zone but dips at a flatter angle (approximately 60° southeast).

As stated above, the deposit is porphyry-style, with very distinct structural controls. The primary mineralizing event was preceded by development of disseminated magnetite-pyrite. Primary economic sulphide mineralization occurred in association with potassic alteration (principally K-spar) resulting in destruction of the earlier magnetite-pyrite assemblage. Following this, carbonate veinlets (principally ankerite) developed along and surrounding the faults, along with varying amounts of pyrite, clay gouge, and sericite.

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10 EXPLORATION

Prior to New Gold’s involvement, the only work that had been done on the deep resource at Afton was the seven holes drilled by Afton Mines in 1973 (two holes) and 1980 (five holes). New Gold commenced exploration work in 2000, with mapping and sampling of the pit (as well as any available outcrop surrounding the pit) and drilling of 96 surface diamond core holes totalling 42,450 m.

In November 2004, an adit was collared in the former Afton open pit at 512 m ASL. A ramp was driven from this portal to provide access for underground sampling, infill drilling to confirm the Mineral Resources, and further exploration drilling to determine the full extent of the mineralization. Total drifting was 2,200 m.

Diamond drilling from February 2005 to April 2006 (cut-off for the 2006 resource estimate) totalled 30,778 m in 78 holes. This brought the total amount of drilling on the Project to 84,755 in 182 holes. Drilling continued throughout 2006, and into 2007, on the C Zone, a deep extension to the New Afton ore body. Five exploration holes totalling 2,996 m were drilled in the Pothook Pit area, which is approximately 0.5 km east of the Afton mine pit. Drilling was also conducted on the Hanging Wall Zones, located under the east side of the Afton pit, in 2008. To date, 276 exploration diamond drill holes have been completed in the entire Project area, for an aggregate length of 130,439 m.

In late 2005, New Gold contracted Fugro Airborne Surveys to complete 1,323 line-km of airborne electromagnetic surveying (DIGHEM) of the Afton and Ajax claims. Quantec Geoscience Ltd. was retained in 2008 to carry out Tensor Magnetotelluric (MT) and DC Resistivity and Induced Polarization (DC/IP) surveys totalling 34.5 line-km.

In addition to the work at Afton, New Gold carried out relatively small programs on the Ajax Group, which is located just over 7 km to the southeast of the Afton pit. Drilling on the Ajax in 2004 and 2006 totalled 4,635 m in ten holes. During the period from 2006 to 2008, New Gold drilled another 16 holes totalling 6,820 m. A radiometric survey, consisting of 44.73 line-km, was carried out in 2007, along with collection of 2,040 geochemical soil samples.

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11 DRILLING

All drilling completed to date on the Project has been diamond core drilling, conducted in a series of programs over a period spanning 2000 to 2008. Table 11-1 summarizes the various campaigns completed over those years. Most of the drilling has been for exploration and resource definition. However, in recent years (2007 and later), geotechnical holes have been drilled for groundwater monitoring and collection of data for mine design.

TABLE 11-1 DRILL SUMMARY BY CAMPAIGN
New Gold Inc. – New Afton Project

Years	Operator	Holes	Size	Metres
2000 to 2003	Atlas	93	NQ and BQ	47,129.15
2005	Boisvenu	67	NQ2	26,516.34
2006	Atlas	14	NQ2	11,152.42
	Boisvenu (U/G)	37	NQ2	17,318.61
2007	Atlas (Surface)	24	NQ	11,137.61
	Atlas (U/G)	5	NQ	1,154.90
	Connors/Foraco*	6	HQ, PQ	5,109.60
	Western	4	NQ, NQ2	1,877.02
2008	Atlas (Surface)	9	HQ, NQ2	3,501.52
	Atlas (U/G) †	12	NQ	3,192.81
	Connors/Foraco‡	1	NQ2	1,617.55
	Western	5	HQ	731.61

* includes wedge holes AF07-125A and AF07-125B (only metreage from wedge to end of hole counted as metres drilled)

† made up of 12 geotechnical holes

‡ includes wedge hole AF07-139a (only metreage from wedge to end of hole counted as metres drilled)

The individual drilling campaigns are described in the following subsections.

2000–2003 PROGRAM

New Gold used Atlas Drilling Company (Atlas) of Kamloops, British Columbia, to do the surface drilling. Atlas employed two diamond drills, a Longyear 38 and Longyear 44, to complete the drilling programs. All of the surface drilling done to date has been with NQ-size equipment (4.76 cm dia.), except for four drill holes which were BQ (3.64 cm dia.).

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A total of 93 diamond drill holes were collared for an aggregate length of 47,129 m (Figure 11-1). All surface drill collars were surveyed by transit and Brunton compass. Drill hole orientations in 2000 beyond hole 2K-12 were measured using a Pajari Bore Hole Survey Instrument, and in 2001 and 2002 a Reflex Easy-Shot Survey Instrument, which records dip and azimuth, was used. Drilling is discussed in further detail in reports prepared by Behre Dolbear (2003, 2004).

At the completion of the 2000 drilling program, the Main Zone was estimated to be 365 m long, 76 m wide (average), and 300 m in vertical extent.

In 2001, ten diamond drill holes extended the length of the Main Zone 280 m to the southwest of the previous drilling. The 2002 infill drilling intersected the Main Zone from the west and provided exploration data on the areas between previously drilled holes. Drill holes extended the measured length of the Main Zone 120 m to the southwest. Drilling in 2003 focused on upgrading Inferred Resources to Measured and Indicated Resources. At that time, the Main Zone was estimated to be 800 m long, 90 m wide (average), and 300 m in vertical extent.

2005–2006 PROGRAM

The 2005 underground drilling program was carried out by Boisvenu Drilling Ltd. of Vancouver, British Columbia, using NQ2 equipment (5.06 cm dia.). All drill holes were surveyed using a photo-bore single-shot instrument for dip and azimuth changes downhole, with collar surveys done by transit.

Underground drilling totalled 26,516 m in 67 holes during 2005, with an additional 37 holes drilled in 2006 (17,319 m). The 2005 program focused on definition drilling of the known mineralization in the Main Zone. The pattern was configured to fill in sections halfway between the existing surface holes, bringing the nominal section spacing to 40 m. In 2006, the principal target area was the C Zone. This drilling was successful in delineating a modest Inferred Mineral Resource of 7.9 Mt grading 0.96% Cu, 0.88 g/t Au, and 1.6 g/t Ag.

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The core from the drill was brought by the drillers on a daily basis to the core shack where it was placed on a pallet and covered until placed in the temporary racks. The core was logged for rock type, mineralization, and alteration, and also for geotechnical data including percent recovery, rock quality designation (RQD), character and number of fractures. The core was photographed prior to being sampled. Scott Wilson RPA noted in the site visits conducted in 2005 and 2006 that the core recovery was generally excellent.

Underground drilling was carried out on 12 sections and consisted of several drill hole fans from each station (Figure 11-1). Drilling completed to that time had delineated a mineralized body measuring about 1,000 m in length, 50 m to 100 m in width, and 400 m in vertical extent, with a moderate plunge to the west.

2007 PROGRAM

The 2007 drilling also targeted the Hanging Wall Zones, as well as depth extensions to the C Zone. Geotechnical drilling was also conducted from both surface and underground. Four holes were drilled on the C Zone, and 23 holes were drilled on Hanging Wall Zones targets. The geotechnical program comprised five underground and four surface holes.

The C Zone program was carried out by FORACO Drilling Ltd. (FORACO). Core sizes were PQ (8.51 cm dia.), HQ (6.35 cm dia.), NQ2, and BQTK (4.07 cm dia.). Three contractors: FORACO, Atlas, and Western Exploration Diamond Drilling Ltd. (Western), were employed for the Hanging Wall program. The core sizes were HQ and NQ2. Atlas and Western were also contracted to do the geotechnical drilling, which was entirely NQ2 size.

Core-handling protocols in 2007 were similar to those applied in 2005-2006. Downhole surveys were carried out using either a Reflex Easy-Shot or Maxibor instrument. The Reflex measurements were made every 50 m for holes greater than 100 m depth. For shorter holes, measurements were made at the mid-point and bottom of the hole. The Maxibor provided readings every three metres downhole.

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2008 PROGRAM

Drilling continued on the Hanging Wall Zones in 2008, with a total of 4,450 m drilled in 15 surface holes. Drilling contractors used were Atlas, Western, and Connors Drilling Ltd.

New Gold now has a formal protocol for locating, monitoring, surveying, and logging the drill holes. Planned holes are plotted in plan and section to confirm the target and collar location. A geologist inspects the collar site, and the surveyors mark both the collar location and line. Once the drill has been set up, the geologist conducts an inspection, checks the orientation of the rig, and confirms with the driller. Daily site inspections are conducted by a geologist as the hole is being drilled.

Following completion, a downhole survey is carried out by a technician using the Reflex instrument (as described under 2007 Program). The Maxibor is no longer used as New Gold personnel have concerns regarding its mechanical reliability.

At this time, also, a request is made to the surveyors for collar pick-up. Scott Wilson RPA notes that, in practice, some underground collars are not actually surveyed, but are located relative to the drift survey. Site reclamation of both the pad and the sump is then carried out.

Core is delivered at the end of each shift to the logging facility housed in a Quonset hut on site. Depending on the back-log, the core is either laid out for processing or stored for later logging. Processed core is laid out, washed, checked for correct depth measurements, and marked with start and end measurements. Any incorrect depth blocks are corrected at this time. Aluminum tags, denoting hole and box numbers, are affixed to each box. The core is then marked for sampling on two-metre intervals.

Geotechnical logging is then carried out, followed by the geological logging. Geological logging includes lithology, alteration, and mineralization. The logs are transcribed into spreadsheets, which are then sent to the database manager along with the core imagery and sample lists.

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Sample tags are prepared for the marked intervals as well as the quality assurance/quality control (QA/QC) samples, and this is all recorded in a spreadsheet. The core is then photographed and placed in racks to await sampling.

In Scott Wilson RPA’s opinion, the core handling protocols are consistent with common industry practice and are appropriate for the Project.

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12 SAMPLING METHOD AND APPROACH

2000-2003 DRILL PROGRAMS

New Gold consultant, James McDougall, P.Eng., reviewed and advised on diamond drilling and geological fieldwork for the 2000 to 2003 programs. All significantly mineralized drill core from the 2001 to 2003 programs (except for specimen sections) was logged, photographed, and diamond sawed, with half the core retained in the core box at New Gold’s core shack and the other half taken directly to the laboratory. Samples were conveyed by laboratory employees from site to Eco Tech Laboratory Ltd. of Kamloops, British Columbia, for analysis for Cu, Au, Ag, and Pd. Core from the 2000 program was split rather than diamond sawed and was not photographed prior to splitting. However, in view of the extreme fragmentation involved with core splitting, occasional one-foot unsplit mineralized sections were retained for structural studies.

Sample lengths were predominantly 2.00 m or 3.05 m, however, Scott Wilson RPA notes that there are a significant number of longer intervals. The maximum sample length recorded in the database is 15.2 m, and 930 samples measure greater than 3.05 m. Approximately 4.7% of the samples contained within the deposit are greater than 3.05 m in length.

A limited number of check samples were randomly selected and sent to Acme Analytical Laboratories Ltd. and International Plasma Laboratory Ltd. of Vancouver, British Columbia. All of the check samples correlated well with the originals (Behre Dolbear 2003). Selected core samples were examined microscopically in the field, and in thin section by petrographer J.F. Harris, Ph.D.

2005 UNDERGROUND SAMPLING

The underground samples comprise continuous chip samples taken on both sides of the drift and correspond with each round of approximately 4.8 m in length. Muck samples are also taken for each round. These samples were not used for mineral resource estimation.

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2005-06 DRILL PROGRAMS

Drill holes were sampled in two-metre lengths or less as marked by the geologist. Scott Wilson RPA notes that, while the sampling protocol includes provision for sampling at less than two-metre lengths, virtually all of the sampling done since the start of 2005 was done on two-metre intervals. Sampling is reportedly based on visible mineralization, but, in practice, most of the core was sampled. The samples were sawn in half, bagged, and put into plastic pails for shipping. The sampler filled out sample tags and, for every 20 samples, included a blank sample prepared from previously drilled barren andesite, a tag for a standard to be put into the sample stream at the lab, and, in addition, a quarter sample to provide a duplicate.

The pails were kept in a locked trailer, which housed the cutting facility, until shipped. The laboratory picked up the samples several times a week. Sample tags were stapled onto the core boxes at the start of each sample. After the core was sampled, the remaining half core was stored in the original boxes on pallets outside. Each pallet was covered with plastic wrap and securely fastened. Scott Wilson RPA notes that a permanent core storage facility with covered racks has since been established on site and all core has been relocated to this facility.

The assay results were loaded directly from the laboratory into an MS Access database. The drill logs were entered into MS Excel spreadsheets from the handwritten logs.

2007 AND SUBSEQUENT DRILLING PROGRAMS

The two metre sampling interval standard has been maintained and is specified in the formal protocol that New Gold now maintains. Samples are marked as described above in Section 11 of this report. Technicians retrieve the boxes from the rack and convey them to the cutting shack. Core is sawn in half, and the sample is placed in a plastic bag marked with the tag number. The remaining half core is put back in the box for storage. Each day, the samples are placed in rice bags, which are then marked with the sample sequence and prepared for shipping. A daily manifest is prepared, detailing the sample numbers cut and packaged, and is provided to the laboratory. Samples are stored in a locked enclosure until pick-up by laboratory personnel. Each shipment is accompanied by the manifest, with a duplicate copy retained at site.

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Starting in early 2007, New Gold commenced a program of assaying pulps by inductively coupled plasma (ICP) in order to build a database of potentially deleterious elements. Starting in the walls, and progressing through the deposit to the opposite wall, every fifth sample was selected from the drill database. These samples were run using ICP. At the time of the site visit, 9,137 samples had been analyzed in this program.

In Scott Wilson RPA’s opinion, the sampling method is appropriate for the deposit type and is consistent with current industry practice. The samples are taken in such a way as to be representative of the deposit as it is presently known.

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13 SAMPLE PREPARATION, ANALYSES AND SECURITY

2000-2003 DRILL PROGRAMS

Sample preparation and analysis were as follows:

·	All samples were sorted, documented, dried (if necessary), roll crushed to -10 mesh, split into 250 g subsamples, and pulverized to 95% -140 mesh.

·	Samples for Cu metallics assay (when requested) were split and pulverized into additional 250 g subsamples of -10 mesh material. The entire pulp was screened to -140 mesh.

·	Au and Pd were subsampled to 30 g aliquots and analyzed with conventional fire assay using atomic absorption (AA) and/or Inductively Coupled Plasma (ICP) finish. Minimum reported detection for Au and Pd was 0.005 g/t.

·	Cu was determined by AA using aqua regia digestion. “Metallic” Cu (when required) included two Cu assays per sample.

·	Ag geochemical analysis was by aqua regia digestion and AA.

·	All equipment was flushed with barren material and blasted with compressed air between each sampling procedure.

All core drilled by New Gold in the 2000-2003 programs was assayed by Eco Tech Laboratories Limited (Eco Tech) of Kamloops, British Columbia, who are Certified Assayers and participate in the National Canmet Proficiency Testing, and who maintain their own in-house QA/QC program.

The New Afton property is fenced and gated, and reasonably secure. It was reported that after the core was logged and sawn, tied sample bags were locked in New Gold’s field office until picked up by personnel from Eco Tech for transport to their facilities. Drill core was stored in core racks at the locked, secure core shack. Rejects were kept at Eco Tech’s office, and pulps were securely stored at New Gold’s field office.

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2005 AND SUBSEQUENT DRILL PROGRAMS

All analytical work was performed by Eco Tech. Copper and silver assays were determined using standard acid digestion followed by AA. Gold and palladium were determined using fire assay followed by an AA finish. The one-in-five deleterious element samples were run for 30-element ICP.

Mr. Stewart Wallis of Scott Wilson RPA visited the Eco Tech Kamloops facility on June 7, 2005, for an inspection, and discussed the assaying methodology and QA/QC protocols with Jutta Jealouse, President. Internal checks consist of a minimum two repeats, one blank, two re-splits, and two or three reference standards, one for Cu, one for Ag, or one combined Cu/Ag and one for Au/Pd. If native Cu is reported on the sample sheets, a metallic screen analysis is run in addition to the regular assay.

The laboratory was observed to be clean and proper procedures were in place to track each sample. Assay results and internal check results were checked and batches rerun if problems were observed. The laboratory has since been acquired by Stewart Group Ltd., of the United Kingdom. Eco Tech is ISO 9001-2000 accredited.

Bulk density measurements are made at Eco Tech, using a water immersion method. Specimens are sealed in wax, weighed in air and in water, and the density determined by the ratio between the weight of displaced water and the weight of the sample. Correction is made for the volume of wax encasing the specimen. Data for 686 determinations were included in the database provided to Scott Wilson RPA for review.

In Scott Wilson RPA’s opinion, the assay methodologies are standard practices in the industry and are appropriate for the mineralization at New Afton. The laboratory is well-known, accredited commercial assayer, with considerable experience with Afton ores. As noted above, the samples are in the custody of New Gold personnel or authorized designates from the drill to the lab, and Scott Wilson RPA has no concerns regarding sample security.

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14 DATA VERIFICATION

2000-2003 DRILL PROGRAMS

New Gold’s QA/QC program was as follows:

·	Assay standards were routinely used to control assay precision.

·	One in nine pulp samples were re-assayed by Eco Tech.

·	One in 25 reject samples were re-split and re-assayed by Eco Tech, and

·	Pulp samples were randomly selected for duplicate assaying by different laboratories.

Previous reports (Behre Dolbear 2003, 2004) have described validation of the assays from the surface drilling program. Behre Dolbear (2004) concluded that the assay and survey database used for the Afton mineral resource estimation was sufficiently free of error to be adequate for resource estimation.

2005-2006 DRILL PROGRAM

Scott Wilson RPA conducted a review of the QA/QC results for the 2006 mineral resource estimate. Independent samples were not collected as the property had been previously visited and reported on by of a number of independent Professional Geologists and is a former producing mine.

New Gold’s protocols were somewhat modified from the ones used prior to 2005. A blank, standard, or duplicate was entered into the sample stream at a frequency of one every eight samples. A total of 18,508 unprepared samples were submitted to Eco Tech during the 2005-2006 program (Konst, 2006). This included 801 blanks, 793 standards, and 721 duplicates, which, in sum, equates to approximately 12% of all samples run.

Sample QA/QC data from the underground drilling program was analyzed by Ron Konst, P.Geo., an independent consultant retained by New Gold. Mr. Konst noted that there were 53 blank assays and 18 standard assays with results outside of an acceptable error limit. These samples comprised 7% and 2% of the total blank and standard assays, respectively, although it was reported that several of these “outliers” were the result of improper labelling (i.e., standards sent as blanks, and vice versa). The Konst report

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recommended that the batches with out-of-specification (OOS) QA/QC data be investigated and re-assayed, if appropriate. This work was carried out and no material changes to the assay database resulted. Several of the OOS blanks were found to be misidentified in the database and did not represent improper assay results.

The duplicate data were analyzed using scatter diagrams and Thompson-Howarth plots to determine if any biases were present and to define the assay precision. The precision for Cu at a 0.6% grade was ±9%, and for Au at a 0.5 g/t grade, ±20%. No biases were detected.

Scott Wilson RPA reviewed the Konst report and the QA/QC data, and conducted independent analyses of the QA/QC assays. The duplicate analyses were subject to Thompson-Howarth plots to confirm the results of the Konst report. Also, each of the blank assays was plotted against the sample that immediately preceded it in the sample stream, to check for evidence of contamination in sample preparation. No concerns were indicated by this analysis. Scott Wilson RPA reviewed the plots of the standard assays and confirmed that there was no indication of systematic bias, nor any trends or particular time periods where assays tended to be OOS.

Scott Wilson RPA concurred with the principal conclusion of the Konst report and considered the assay data suitable for use in mineral resource estimation. However, in the course of preparing the estimate, Scott Wilson RPA found a number of database errors, which were brought to the attention of New Gold personnel and corrected. These errors were often observed to persist in subsequent versions of the database, indicating that the validation process was not robust. Eventually, these errors were expunged, but there were concerns regarding the data handling.

Scott Wilson RPA carried out a check of approximately 10% of the drilling database against the original assay and survey records. Drill holes were selected on a more or less random basis, spanning the entire life of the Project to date. No errors were found.

2007 AND SUBSEQUENT PROGRAMS

New Gold consultant, Bruce Davis, reviewed the QA/QC data in October 2008 and found that QA/QC results were showing significant deficiencies but remedial measures had not

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been taken. The recommendation was made to select approximately 300 sample pulps for re-assay of Cu and Au, both at the principal laboratory, Eco Tech, as well as at another laboratory. The results of this re-assay program did not indicate that there were any concerns with respect to bias. However, the data did display a somewhat high level of scatter for pulp duplicates.

Scott Wilson RPA selected approximately 10% of the holes drilled since the last resource estimate (2006). Records showed that 110 holes had been drilled and, of these, 11 were checked against the original assay data received from the laboratory. No errors were found. In addition, the entire database was checked using the Gemcom validation utility. One or two very minor discrepancies were found with respect to the header information, but none that would impact on resource estimation.

Present protocols are for the assay data to be sent via email and hard copy to the database manager at site. These data are then converted to comma-delimited files and imported into the master database for validation. The validated data can then be exported to various users for specific applications such as geological interpretation, plotting, and resource estimation.

At the time of the most recent Scott Wilson RPA site visit, a new database system was being implemented, with automated functions for importing, validating, and exporting the drill data. Data handling protocols were also being introduced that would provide for the establishment of a single master database, with one person responsible for its maintenance. Scott Wilson RPA considers this to be a very worthwhile step and strongly recommends that all phases of the data handling process be rigorously monitored. Assay QA/QC results should be reviewed on a continual basis, in a timely fashion, and remedial actions taken when necessary.

In Scott Wilson RPA’s opinion, the present protocols for data acquisition and management appear to be reasonable.

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15 ADJACENT PROPERTIES

The New Afton Project is the most advanced exploration and development project in the area. The ground adjacent to the southeast (Ajax deposit) is 100% owned by Abacus Mining & Exploration Corp. In July 2009, Wardrop Engineering Inc. produced a Preliminary Assessment on the property which contemplated a 60,000 tpd open pit mine and mill facility. Measured and Indicated Mineral Resources for the project were quoted as 422.4 Mt at a grade of 0.30% Cu, 0.19 g/t Au, 0.0015% Mo, and 0.31 g/t Ag, with an additional Inferred Resource of 80.6 Mt grading 0.22% Cu, 0.16 g/t Au, 0.0011% Mo, and 0.38 g/t Ag (Table 15-1). The database for the 2009 Mineral Resource estimate comprised 140,490 m of diamond drilling in 366 drill holes along with 53 percussion drill holes totalling 4,386 m.

TABLE 15-1 AJAX DEPOSIT – MINERAL RESOURCE ESTIMATE
Abacus Mining & Exploration Corp. – Ajax Deposit

	Cut-off CuEq* (%)	Mt	Cu (%)	Au (g/t)	Mo (%)	Ag (g/t)	CuEq*(%)
Measured	0.13	231.4	0.30	0.18	0.0018	0.35	0.32
Indicated	0.13	211.0	0.29	0.19	0.0012	0.31	0.32
Measured + Indicated	0.13	422.4	0.30	0.19	0.0015	0.31	0.32
Inferred	0.13	80.6	0.22	0.16	0.0011	0.38	0.24

*A recovered copper equivalent grade (CuEq %) is calculated using copper and gold commodity prices of US$2.00/lb and US$700/troy oz (respectively) in the formula:

CuEq (%) = [Cu(%) x Cu-Rec x 22.0462 x $2.00 + Au(g/t) x Au-Rec x 0.03216 x $700]

[22.0462 x $2.00]

Revised metal recoveries were used, which were developed during the preparation of the PA:

Cu-Rec = 32.591 x Cu(%) + 72.732, and Au-Rec = 33.87 1 x Au(g/t) + 75.29 (max. value = 90%).

Scott Wilson RPA has compiled information from reports filed on SEDAR. Scott Wilson RPA has been unable to verify information on adjacent properties, and this information is not necessarily indicative of the mineralization on the property that is the subject of this Technical Report.

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16 MINERAL PROCESSING AND METALLURGICAL TESTING

This section has been summarized from the Hatch 2007 Feasibility Study as well as the SGS Mineral Services (SGS) metallurgical test report, both completed in April 2007. Where necessary, the information on equipment sizing has been updated with recent data provided by New Gold on the implementation phase of the Project.

METALLURGICAL TESTING

A metallurgical test program was developed by Hatch in conjunction with New Gold and SGS. The primary test work was completed by SGS from March to October 2006, while secondary programs were completed by Pocock Industrial Inc. (Pocock) and Knelson. Mineralogy was completed by Advanced Mineral Technology Laboratory (Amtel) and Vancouver Petrographics Ltd. (VPL). Hatch reviewed and interpreted the results from the various test programs to develop the process design criteria and flowsheets.

Three ore types represent the mineralization for this copper-gold deposit. At the testing stage, the proportion of these ore types was estimated based on the resource estimate at the time. The preliminary metallurgical evaluations focused on the mesogene and hypogene ore types. The metals of economic interest were copper, gold, silver, and palladium. Table 16-1 provides a comparison of the estimated percentages for the test program with the percentages in the 2009 Life of Mine (LOM) plan. In the opinion of Scott Wilson RPA, the samples used in testing were reasonably representative of the different ore types.

TABLE 16-1 PERCENTAGE OF ORE TYPE
New Gold Inc. – New Afton Project

Rock Type	% Estimated for Testing	% in LOM Plan
Hypogene (Primary Zone)	50%	52%
Mesogene (Transition Zone)	40% to 50%	44%
Supergene (Enriched Zone)	5% to 10%	4%

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PROGRAM DESCRIPTION

This metallurgical test program was conducted in two phases. The first phase was conducted on bulk samples representing the mesogene and hypogene ore types, to establish typical metallurgical requirements for the two main ore types. These samples were obtained from the exploration decline into the deposit. Barren hanging wall and footwall materials adjacent to the ore were obtained to evaluate the impact of dilution.

The second phase evaluated the variability of metallurgical response to the standard conditions established. This phase of the program employed split drill core samples from various spatial locations in the orebody. In addition, split drill core samples from diamond drill holes 90 and 91 were included. Drill hole 90 represented two different supergene ore types, while drill hole 91 was selected to more thoroughly represent the mesogene ore. SGS completed grinding and flotation tests on all of these samples and also conducted solid-liquid settling tests on flotation concentrates.

Pocock conducted tailings dewatering tests on various rougher flotation tailings samples. Their scope of work included flocculant screening, static and dynamic gravity sedimentation, and pulp rheology, pressure and vacuum filtration tests.

Amtel completed a mineralogical evaluation and reported on the deportment of gold values in the three ore types. VPL issued two mineralogical examination reports, one on 30 bulk samples for the two main ore types and the second on 70 core and bulk samples representing the three ore types, footwall, and other rock units.

SAMPLES

In the first phase, bulk samples were taken to represent the two main ore types, the mesogene and the hypogene. For each ore type, three grades were prepared.

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TABLE 16-2 BULK SAMPLES
New Gold Inc. – New Afton Project

Rock Type	Low Grade		Medium Grade		High Grade		Average
	%Cu	g/t Au	%Cu	g/t Au	%Cu	g/t Au	%Cu	g/t Au
Hypogene	0.82	0.34	1.45	1.06	2.19	0.91	1.49	0.77
Mesogene	0.75	0.68	1.56	1.13	1.48	0.54	1.26	0.78

For each bulk sample a total of nine 300 kg drums were collected and shipped to SGS for metallurgical testing, three drums for each grade range. Composite mixing proportions were then calculated to blend the three grade samples from each ore type to match a target assay. The final mesogene master composite assayed 1.44% Cu and 1.26 g/t Au, while the hypogene master composite assayed 1.26% Cu and 1.05 g/t Au. Scott Wilson RPA notes that the LOM plan grades are lower at 0.96% Cu and 0.72 g/t Au, and this may have an effect on the metallurgical performance.

Dilution samples were submitted for ore blending to represent the expected dilution with the selected block caving mining method. Block caving will result in lower selectivity at the ore interfaces. To evaluate this effect on the metallurgical performance, a total of approximately 500 kg barren hanging wall and footwall samples were sent to SGS. The dilution type samples represented the mesogene, hypogene, and picrite rock types.

Supergene samples consisting of split drill core from drill hole 90 were also tested. Six pails, weighing a total of 193 kg, were sent to SGS. Half of the sample was designated Type 1, representing a hematite-rich alteration, and half was Type 2, representing a weaker hematite alteration but with more interval fractures.

For the second phase, variability samples were selected to represent various spatial locations within the mineralized zones. For this work, 18 pails of split drill core samples representing supergene, mesogene, and hypogene ore types and the three grade ranges were submitted. Each pail weighed 24 kg to 30 kg.

In addition nine pails of core samples from drill hole 91, weighing 245 kg, were sent to SGS. These additional core samples were submitted to provide an expected range of primary and secondary copper minerals representing all three ore types.

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MINERALOGY

Three laboratories conducted mineralogical studies on the New Afton samples.

•	VPL conducted two mineralogical studies: one for hypogene and mesogene ores on the underground bulk samples and a second petrographic study on 70 drill core and underground samples for all three ore types and footwall rocks.

•	SGS conducted a modal analysis on the mesogene and hypogene ores and a mineralogical study on hypogene copper concentrate samples.

•	Amtel conducted a gold deportment study for all ore types.

The VPL studies are briefly summarized below:

•

Hypogene ore contains chalcopyrite (CuFeS2) as the main copper mineral of interest with a chalcopyrite:pyrite ratio of 25:1. Chalcopyrite occurs primarily as free grains but with inclusions in silicates. Gold particles varied in size from 5 µm to 40 µm and occur as native gold between pyrite and chalcopyrite grains. There is sericite present.

•

Mesogene ore copper mineralization includes chalcopyrite, as well as bornite (Cu5FeS) and chalcocite (Cu2S) in varying proportions that occur as intergrowths replacing the original chalcopyrite mineralization. Chalcocite tends to be finer grained and dispersed, which will have a negative influence on the final concentrate grade. Pyrite is extremely rare. The content of sericite is higher. No gold particles were observed.

•	Supergene ore copper minerals are mainly chalcocite and native copper. Native copper occurs as micro-fracture coatings, which will render recovery by flotation difficult. Lower levels of sericite and carbonate compared to the mesogene ore were observed.

•	Footwall Zone rocks resemble mesogene rocks but with lower levels of sericite. Some sulphides are present as barren pyrite, but copper minerals are rare.

The SGS modal analysis quantified the levels of various minerals. This study identified tennantite (Cu11FeAs4S13) as the primary source of arsenic (As). The As level in mesogene ore poses an economic concern for the concentrate produced from this ore type. The study also indicated the presence of tetrahedrite (4Cu2S·Sb2S3) in the mesogene ore type, which has the potential to cause higher Sb levels in the copper concentrate, but this was not evident in flotation test results.

The levels of sericite and chlorite were high enough in both the mesogene and hypogene rock types to cause slimes problem in the dewatering of flotation tailings.

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Subsequently, it was concluded that tailings would not be thickened as part of the process (see Mineral Processing section).

The Amtel study noted that approximately 40% of the gold was associated with copper sulphides and 30% occurred as liberated electrum in all three ore types.

GRINDING

A detailed study of the grinding characteristics for the mesogene and hypogene ore types was undertaken. Coarse ore was subjected to energy determinations using the Bond Low-energy Impact Test and the JKTech Drop-weight Test (DWT).

Coarse ore was crushed and homogenized to form two grindability composite at minus 1-1/4” and minus six mesh. These samples underwent further grinding tests including the SAG Power Index (SPI) test, the MacPherson 18” Autogeneous grinding test (AWI), the Bond Rod Mill work index (RWI), the Bond Ball Mill work index (BWI), and Abrasion Index (AI) determinations. Those results have been summarized in Table 16-3.

TABLE 16-3 GRINDING INDEX SUMMARY
New Gold Inc. – New Afton Project

Rock Type	Dry Solids S.G.	SPI (mins)	AWI (kWh/t)	RWI (kWh/t)	BWI (kWh/t)	AI (g)
Hypogene	2.70	200	19.2	18.5	21.8	0.164
Mesogene	2.77	140	18.2	18.3	19.8	0.071

These work indices have classified both ores as being hard and competent, with the hypogene ore being very hard. For the Abrasion index classifications, the mesogene ore has been categorized as low-abrasive and the hypogene ores as mildly abrasive.

FLOTATION

A total of 83 batch flotation tests and seven locked cycle tests (LCT) were performed over the entire test program. The first phase test work on the bulk samples included open circuit and locked cycle tests and evaluated the following:

·	primary and regrind product sizes

·	collector and frother types

·	flotation kinetics

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The effect of dilution with barren rock on the metallurgical performance was studied. A limited amount of flotation work was conducted on the supergene ore. At the end of the test program, a flowsheet was established that consisted of two stages of primary grinding, rougher flotation, rougher concentrate regrinding, first stage cleaning followed by cleaner-scavenger flotation, and two additional stages of cleaner flotation. The cleaner-scavenger concentrate was recycled for regrinding. Rougher tailings and the cleaner-scavenger tailings are handled separately but combine to make final tailings.

The main flotation reagents used were collectors, including potassium amyl xanthate (PAX), sodium isopropyl xanthate (SIPX), Aero 5100, Aero 7249 and Cytec 3418A (a phosphine reagent); and frothers, including methyl isobutyl carbinol (MIBC), Dowfroth-250 (DF250) and pine oil. PAX was selected as the collector based on its traditional role in porphyry flotation, while DF250 was used as the principal frother. The high sericite contents of some of the ores will warrant the possible use of carboxyl-methyl-cellulose (CMC). Subsequent testing determined the ultimate selection of PAX, 3418A, MIBC, and lime as the reagent suite for the process plant.

The results indicated that a P80 of 160 µm was the optimum primary grind. Finer primary grinding of the hypogene ore to a P80 of 106 µm did not improve the copper recovery. Gold flotation kinetics followed that for copper for both ore types.

The optimum regrind size for cleaner flotation performance was between 21 µm and 59 µm for the hypogene and mesogene ores. Consequently, a P80 of 35 µm was selected for sizing the regrind mill.

Dilution tests on bulk samples included the addition of 10% and 20% hanging wall and footwall materials for the hypogene (five tests) and mesogene (eight tests) ore types. This was followed by a series of four cleaner tests on mesogene ore, diluted with supergene ore by 20% and 40%. The results indicated that the diluting rock produced similar rougher copper and gold recovery but with higher mass recoveries and subsequently lower concentrate grades. The impact of this would be an increased load on the regrind mill and cleaner circuit.

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LOCKED CYCLE TESTS

Locked cycle tests were conducted on bulk samples in duplicate for the main ore types, with a different frother reagent (MIBC or DF250) used on each. The results in Table 16-4 show that MIBC provides a higher concentrate grade but a lower copper recovery. The hypogene ore exhibits slightly better metallurgical recovery than the mesogene ore.

	TABLE 16-4 LOCKED CYCLE TEST SUMMARY
New Gold Inc. – New Afton Project

Test	Combined	Weight	Assays		Recovery
No.	Product	(%)	Cu %	Au g/t	Cu %	Au %
LCT-H1	Cu 3rd Cl Conc	3.9	29.4	15.3	91.6	84.9
Hypogene	Combined Final Tail	96.1	0.10	0.10	7.7	14.3
MIBC	Calc Head	100.0	1.25	0.71

LCT-H2	Cu 3rd Cl Conc	4.9	24.5	14.3	93.8	93.0
Hypogene	Combined Final Tail	94.1	0.06	0.04	6.0	6.9
DF250	Calc Head	100.0	1.28	0.75

LCT-M1	Cu 3rd Cl Conc	4.0	28.6	17.8	88.6	79.8
Mesogene	Combined Final Tail	96.0	0.185	0.20	13.4	20.5
MIBC	Calc Head	100.0	1.33	0.92

LCT-M2	Cu 3rd Cl Conc	4.8	25.2	16.2	89.4	84.4
Mesogene	Combined Final Tail	95.2	0.139	0.15	9.8	15.4
DF250	Calc Head	100.0	1.36	0.93

A full analysis of the concentrate from the locked cycle tests is provided in Table 16-5, including the results from a variability test on ore from drill hole 91.

TABLE 16-5 CONCENTRATE ASSAYS

New Gold Inc. – New Afton Project

Test		LCT-M1	LCT-H1	LCT-H91-A
Element	Unit	Mesogene	Hypogene	Mesogene
Cu	%	29.7	30.5	30.0
Au	g/t	17.5	15.9	30.2
Ag	g/t	45.9	40.0	75.7
Pd	g/t	2.55	0.25	2.55
Pt	g/t	0.18	0.10	0.10
Pb	%	<0.01	<0.01	0.02
Zn	%	0.12	<0.01	0.20
Fe	%	24.3	28.8	13.0

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Test		LCT-M1	LCT-H1	LCT-H91-A
Element	Unit	Mesogene	Hypogene	Mesogene
ST	%	27.9	30.7	17.3
As	%	0.83	<0.01	0.78
Sb	%	0.11	<0.01	0.14
Se	g/t	43	<40	85
Hg	g/t	33.6	0.6	87.2
Cl	g/t	58	55	N/A
F	g/t	200	100	N/A
Insoluble	%	12.1	7.1	29.6

There will be gold, silver, and likely palladium credits in selling the concentrate. Smelter penalties will be incurred for arsenic and mercury when treating a significant portion of mesogene ore, and possibly for antimony as well, as typical threshold values are 0.2% As, 0.1% Sb, and 10 ppm Hg.

The insoluble values were low for the first phase locked cycle tests but they were high in the drill hole 91 concentrates. Higher insoluble values would have been expected due to the high clay content of these ore types. Insoluble minerals can produce more difficulties in flotation and an insoluble depressant or dispersant may be required in the flotation reagent suite.

After the locked cycle tests were completed, separate reagent suites were evaluated to determine whether improvements could be made to the flotation response for various ore types. As the block caving mining method was planned to be used, the focus was on the development of a single flowsheet for all ore types.

GRAVITY TESTS

Gravity test work was also completed on hypogene, mesogene, and mesogene-supergene blends to optimize recovery of gold to the final concentrate. The gravity concentration work by Knelson identified the presence of gravity recoverable gold. An economic benefit for gravity concentration was not identified from the test work, but is based on good experience at other porphyry copper-gold properties. Gravity concentration circuits will be included in the primary and regrind circuits.

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Subsequent test work determined that final grade concentrate could be produced from coarse material with flash flotation in the grinding circuit. A flash flotation circuit was subsequently added to the process plant design.

In addition, Knelson conducted a preliminary bench-scale assessment of native copper recovery by gravity concentration for the three ore types. This was done because the native copper values proved to be difficult to recover by flotation methods. Based on the analysis completed by Knelson, there will be potential to recover some of the copper in the supergene ore to a gravity concentrate at a marketable concentrate grade. Copper in hypogene and mesogene ores did not appear to be amenable to gravity concentration. More work on a larger scale should be completed to quantify the native copper impact on metallurgy.

VARIABILITY TESTS

In the second phase of flotation work, tests were conducted on drill core samples to determine the variability of flotation response on a spatial basis throughout the different zones of mineralization. Various intervals from sixteen drill holes were provided for this evaluation. Within these intervals, the samples were classified by ore type and by copper grades, as was done in the first phase. A total of 25 open circuit batch cleaner tests were performed, ten on hypogene ore, seven on mesogene ore, and eight on supergene ore. A summary of the variability test results is shown in Table 16-6.

Later in the test program, another set of variability samples from drill hole 91 were provided as a more wide representation of the mesogene ore type; those samples were provided as three composites, labelled A, B, and C. The sample designations were a function of the depth in that drill hole. The “A” composite is from the 266 m to 300 m depth, the “B” composite from 300 m to 360 m depth, and the “C” composite from the 360 m to 418 m depth. Testing on these samples included five batch tests and three locked cycle tests.

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TABLE 16-6 VARIABILITY (BATCH) TEST SUMMARY RESULTS
	New Gold Inc. – New Afton Project

Sample	Calculated Head		Concentrate Grade	Concentrate Recovery
	Cu, %	Au, g/t	Cu, %	Cu	Au
Hypogene
Average	1.28	0.95	27.8	87.8	86.0
Max	2.06	1.31	29.9	90.4	90.6
Min	0.55	0.50	25.2	85.3	80.7
Mesogene
Average	1.50	0.60	33.1	75.5	72.0
Max	2.78	1.06	45.7	88.8	88.2
Min	0.63	0.18	27.7	60.2	51.6
Supergene
Average	1.65	0.62	44.1	81.8	74.8
Max	2.47	1.38	55.5	87.6	85.0
Min	0.96	0.17	33.2	69.5	57.4

Metallurgical recoveries have been analyzed by New Gold and Hatch and directly related to the head grade and to grind size from the variability tests. The relationships with head grade are shown in detail in the subsection on Metallurgical Recovery. This effect of head grade can be seen in Table 16-6, most notably for mesogene ore, with copper recovery ranging between 60% and 89% depending on head grades ranging from 0.63% Cu to 2.78% Cu. Gold recovery ranges are even more pronounced than copper. Copper concentrate grades produced from all samples were acceptable and deemed to be saleable.

DEWATERING TESTS

Dewatering tests were conducted on copper concentrates and on rougher tailings. SGS conducted static settling tests on final copper concentrates and vacuum filtration tests on rougher tailings samples. Pocock conducted a series of dewatering tests on rougher tailings samples. The Pocock scope of work consisted of flocculant optimizing, gravity sedimentation, pulp rheology, pressure and vacuum filtration tests.

CONCENTRATE DEWATERING TESTS

Concentrate from two bulk flotation tests were subjected to settling tests. The primary and regrind P80 sizes for these tests were 110 µm and 27 µm, respectively. Copper

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concentrate settling test resulted in underflow densities of 61% to 64% solids and unit areas of 0.028 m2/tpd to 0.058 m2/tpd with the addition of flocculant and coagulant. Subsequent testing was done by an equipment supplier to support the selection of the equipment for the process plant.

No pressure filtration tests were included in the test program because of the large quantity of concentrate required for conducting such evaluations. Subsequent testing was done by vendors to support the bidding process and the selection of the equipment for the process plant.

ROUGHER TAILINGS DEWATERING TESTS

Dry stacking of tailings was initially evaluated by SGS with vacuum filtration tests conducted on rougher tailings under various operating conditions. Cake moistures varied from 18% to 21%, with filtration rates varying between 11 and 79 dry kg/m2·h.

Pocock ran a series of sedimentation tests on the three rock types and for one composite with hanging wall material. Good settling rates and supernatant clarity were obtained using 50 g/t to 80 g/t of a cationic coagulant (Hychem 606) and 20 g/t to 30 g/t of an anionic polyacrylamide (Hychem 303). The result was typical underflow densities of 57% to 64% and unit areas of 0.125 m2/tpd to 0.300 m2/tpd. However, it was concluded that these flocculant and coagulant addition rates were cost-prohibitive and conventional slurry deposition of tailings was adopted for the final design.

METALLURGICAL RECOVERY PROJECTIONS

The data from the variability tests have been analyzed to develop equations for projecting the recovery of copper, gold, silver, and palladium and the deportment of the penalty elements, arsenic and mercury, in the concentrate for all three ore types. Where available, locked cycle results have been used to support the projections from the batch test data points. The equations have been used for projecting production from mining blocks and in the economic model.

COPPER RECOVERY

Copper recovery has been correlated with the copper head grade for a constant copper concentrate grade for all three ore types. Recoveries have been normalized to a 28%

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Cu concentrate grade for hypogene ore, which contains primarily chalcopyrite, to a 30% Cu concentrate grade for mesogene ore, due to the presence of bornite, and 58% Cu concentrate for supergene ore, due to the presence of native copper.

A recovery increase of 2% to 3% has been applied to the best fit equations based on the expected improvement in project plant performance vs. batch open circuit cleaner data. This increase is supported by locked cycle test results for hypogene and mesogene ore types. Typically, locked cycle recovery would be higher than that from batch open cleaners for the same concentrate grade and it would be indicative of the recovery that may be achieved in an optimized plant operation. For supergene ore, the correlation has been developed based on flotation only with an increase to allow for potential native copper recovery by gravity concentration.

For Hypogene ore, the correlation for copper recovery is expressed as:

% Cu recovery = 6.2059·Ln (F) + 90.69

where F = % Cu(ore)

For Mesogene ore, the correlation is based on tails losses and is expressed as:

% Cu recovery = 100(28.062 – 3.168/F)

(29.8944 – 0.0646·F)

where F = % Cu(ore)

For Supergene ore, the correlation is estimated as follows:

% Cu recovery = 41.308·Ln (F) + 54.833

where F = % Cu(ore)

GOLD RECOVERY

Amtel’s mineralogical study concluded that gold is primarily associated with copper sulphides that will be recoverable by flotation in all three ore types. For hypogene ore, the gold recovery did not correlate well with copper recovery, but the gold concentration ratio correlated well with copper concentration ratio. For mesogene and supergene ore types, there was a reasonable direct correlation between gold recovery and copper recovery.

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For Hypogene ore, gold recovery was calculated using the concentration ratio correlation and the calculated concentrate production based on copper:

g/t Au (conc) = 1.0447x - 1.8458

g/t Au (ore)

where x = %Cu(conc)

%Cu(ore)

For Mesogene ore, gold recovery was calculated as follows:

% Au recovery = 15.754·e0.0193x

where x = % Cu recovery

For Supergene ore, gold recovery was calculated as follows:

% Au recovery = 0.6545x + 24.715

where x = % Cu recovery

SILVER RECOVERY

For both hypogene and mesogene ore types, silver recovery has a poor direct correlation with copper recovery, but the silver concentration ratio correlated well with copper concentration, similar to the hypogene gold recovery.

For Hypogene ore, silver recovery was calculated using the concentration ratio correlation and the calculated concentrate production based on copper:

g/t Ag (conc) = 1.181x - 10.271

g/t Ag (ore)

where x = %Cu(conc)

%Cu(ore)

For Mesogene ore, silver recovery was calculated using the concentration ratio correlation and the calculated concentrate production based on copper:

g/t Ag (conc) = 0.7922x – 2.2975

g/t Ag (ore)

where x = %Cu(conc)

%Cu(ore)

For Supergene ore, silver recovery was calculated as follows:

% Ag recovery = 0.714x1.067

where x = % Cu recovery

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ARSENIC AND MERCURY CONTENT IN CONCENTRATE

For hypogene and mesogene ore types, the data showed a strong correlation between the arsenic and mercury concentration ratios and copper concentration ratio. These correlations were used to determine the economical concentrate grade (particularly for mesogene ore which contains higher arsenic) by assessing the effect of concentrate copper grade on the levels of these elements in the concentrate and their associated penalties.

For Hypogene ore, the concentration relationships are as follows:

% As (conc) = 0.8615x - 6.2294

% As (ore)

g/t Hg (conc) = 1.162x - 11.434

g/t Hg (ore)

where x = %Cu(conc)

%Cu(ore)

For Mesogene ore, the concentration relationships are as follows:

% As (conc) = 0.6285x + 6.0855

% As (ore)

g/t Hg (conc) = 0.3864x + 8.2154

g/t Hg (ore)

where x = %Cu(conc)

%Cu(ore)

For Supergene ore, the arsenic and mercury content in the concentrate correlated directly with their respective ratio to copper in the feed as follows:

% As(conc) = 40.658x - 0.0716

where x = %As(ore)

%Cu(ore)

g/t Hg(conc) = 51.077x - 8.4396

where x = g/t Hg(ore)

%Cu(ore)

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MINERAL PROCESSING

The New Afton processing facilities have been designed to process 4 Mtpa of ore and recover copper, gold, and silver. A conventional crushing, grinding, gravity concentration and flotation process has been proposed for the Project, utilizing standard processes and equipment.

The process design criteria were developed based on a combination of test work results, experience from similar operations, and industry practice. The following process description should be read in conjunction with the process flow diagram shown in Figure 16-1.

PROCESS DESCRIPTION

The New Afton mill has been designed to process a blend of primary, transition, and supergene ores. The process will utilize conventional crushing, grinding, and concentration processes to separate valuable minerals from gangue into concentrates at marketable grades. Mineral separation will be achieved by gravity concentration and differential flotation. A regrinding stage will be included in the flotation circuit.

The New Afton mill has been designed to process 11,000 tpd of ore at full capacity and will operate 24 hours per day, 365 days per year. The plant will consist of the following unit operations.

•	Primary crushing

•	Conveying from underground to surface

•	Stockpiles – coarse ore storage and reclaim, and intermediate waste storage

•	SAG mill primary grinding with cone crushing of recycle pebbles

•	Ball mill secondary grinding with closed circuit cyclone classification

•	Gravity concentration

•	Flash Flotation

•	Rougher flotation

•	Rougher concentrate regrinding, classification and gravity concentration

•	Cleaning flotation in three stages with one open stage of cleaner-scavenger

•	Concentrate thickening and storage

•	Concentrate pressure filtration

•	Concentrate storage, loading and weighing (on-site)

•	Tailings cycloning and water reclaim

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PRIMARY CRUSHING, COARSE ORE STORAGE AND RECLAIM

The ore will be crushed underground by a jaw crusher to 150 mm and conveyed to the surface onto a 10,000-tonne live load coarse ore stockpile. Coarse ore will be reclaimed from the stockpile by one of two apron feeders to the SAG mill feed conveyor belt. A dust collector system will be installed in the reclaim tunnel to collect dust generated at the apron feeder.

GRINDING

Coarse ore at 150 mm will be ground in an 8.3 m x 3.96 m SAG mill (7,000 HP) and a 5.5 m x 9.75 m ball mill (7,000 HP) to reduce the ore to a product size of 80% passing 160 µm.

The design criteria for the grinding circuit have been based on hypogene ore, as it has the higher work index. A SAG work index of 19.2 kWh/t and a Ball Mill work index of 21.8 kWh/t have been used with an expected transfer size of 1 mm. In the earlier years, the mill feed will be exclusively hypogene, while later on in the LOM plan there will be various blends of hypogene and mesogene ores.

The SAG mill product will be classified on a 2.44 m x 6.1 m single-deck vibrating screen. The screen oversize will be reduced in a 2,100 mm cone crusher (500 HP) to 12 mm and then recycled to the SAG feed conveyor. The screen undersize will be classified in a bank of six 660 mm diameter cyclones.

The cyclone overflow will feed the rougher flotation cells. The cyclone underflow will be returned to the SAG mill, while a portion of the cyclone underflow slurry will feed the gravity concentrator that will be installed to optimize gold recovery. Gravity concentrates will be directed to the concentrate thickener. Another cyclone underflow stream will be directed to a flash flotation circuit consisting of a 20 m3 rougher and a 5 m3 cleaner to remove coarse copper sulphide particles.

FLOTATION

A single bank of 100 m3 flotation cells is provided for the rougher flotation. Flotation will be conducted at the natural pH, with PAX, 3418A, and MIBC added at the rougher flotation stage. Rougher concentrates will be reground and upgraded to a final

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concentrate grade with three stages of cleaner and one stage of cleaner-scavenger flotation.

Rougher and recycled cleaner-scavenger concentrates will be reground in a 1,250 HP vertical grinding mill operating in closed circuit with three 381 mm diameter cyclones. A second collector, Aerophine 3418A will be added to the regrind mill. The regrind mill will produce a particle size distribution of 80% passing 35 μm. The cyclone overflow will then advance to the cleaning stages.

The first cleaner consists of a bank of seven 20 m3 flotation cells operating in open circuit, while the second and third cleaners will operate in closed circuit. The first cleaner concentrate will be fed to the second cleaner circuit, while the first cleaner-scavenger tails will report to the final tails pump box. The second cleaner concentrate will feed the third cleaner circuit to produce the final copper concentrate. The second and third cleaner circuits consist of 5 m3 flotation cells.

DEWATERING AND CONCENTRATE TRANSPORTATION

The concentrate dewatering circuit includes thickening and pressure filtration. The final concentrates from the third cleaner flotation stage will be dewatered in an 18 m diameter thickener with flocculant added to the solution to enhance solids settling and to achieve the target underflow density of 65% solids by weight. The thickener overflow will be re-used as process water in the plant.

The underflow from the thickener will be pumped to a concentrate stock tank that will be designed for 24-hour storage. Stock tank concentrate will be pumped to one of two 40 m2 pressure filters to produce a concentrate cake containing 8% to 12.5% H2O. The concentrate filter cake will be discharged onto a 1,320-tonne capacity concentrate stockpile in the loading shed. Concentrates will be sampled, loaded into B-Line trucks, and weighed prior to shipment from the site.

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TAILINGS DISPOSAL

The tailings handling system will pump the combined tailings (from the rougher flotation circuit and the cleaner scavenger circuit) to the Cyclone Plant, a set of primary cyclones, located adjacent to the tailings storage facility (TSF).

TAILINGS STORAGE FACILITY

The TSF includes the Pothook pit, the tailings dam, and the impoundment zone. The concept of “dry-stacking” for tailings disposal was evaluated earlier and dropped for various reasons, principally related to the prohibitive costs of the flocculants and coagulants. The plan is to create a centreline raised tailings dam with an upstream (u/s) low-permeability core of compacted glacial till and a downstream (d/s) shell of cycloned sand from the tailings. Certain segments of the tailings dam will also include a starter dam constructed solely out of native glacial till. Cyclone overflow (fines) and the unused whole tailings will be impounded behind the dam, along with a small amount of operating water and seasonal surplus water. The impounded water is to be reclaimed and reused in the plant.

The TSF partially overlies former waste rock dumps and partially is located on native ground. Areas covering waste rock dumps will be lined with a 1.5 mm thick geomembrane to prevent piping into the waste rock.

Initially, whole tailings will be stored in Pothook pit followed by a stage during which a layer of whole tailings will be deposited in the entire impoundment area. Subsequently, a cycloning operation consisting of primary and secondary cyclones will go in operation to separate the sand fraction from the whole tailing.

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The tailings deposition design is based upon the feasibility production schedule which commenced with two years at 1.6 Mtpa before ramping up to 4 Mtpa and contained a LOM plan of 44.4 Mt. The revised production plan contains 47.4 Mt and moves immediately to a production rate of 4 Mtpa. New Gold has modified the tailings dam design and schedule to accommodate these changes.

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17 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

MINERAL RESOURCES

HISTORY

Mr. Gary Giroux, P. Eng., completed an initial resource estimate for the Project in May 2003, which was then updated in January 2004. The estimate was prepared using a block model constrained by a wireframe model of the mineralized body. Grade was interpolated into the blocks using Ordinary Kriging (OK). The January 2004 Measured and Indicated Mineral Resource estimate for the Afton Main Zone at a cut-off grade of 0.7% copper equivalent (CuEq) was 68.7 Mt at a grade of 1.08% Cu, 0.85 g/t Au, 2.62 g/t Ag, and 0.12 g/t Pd. Total contained metal was 1,640 million pounds of Cu and 1.9 million ounces of Au. There was an additional Inferred Mineral Resource of 7.45 Mt at 0.92% Cu, 0.78 g/t Au, 2.34 g/t Ag, 0.12 g/t Pd (Behre Dolbear, 2004).

Following completion of definition drilling, in 2006, Scott Wilson RPA re-estimated the Mineral Resources at New Afton using a block model constrained by wireframe models of the principal mineralized zones. Grades for Ag, Au, Cu, and Pd were interpolated into the blocks using OK. The wireframe models were constructed by New Gold personnel and inspected by Scott Wilson RPA prior to their use in the block model. High grades for all four elements were capped before compositing.

Measured and Indicated Mineral Resources totalled 65.7 Mt grading 1.02% Cu, 0.77 g/t Au, and 2.59 g/t Ag. In addition, a modest Inferred Mineral Resource of 7.9 Mt grading 0.96% Cu, 0.88 g/t Au, and 1.6 g/t Ag was estimated for the C Zone.

INTRODUCTION

The most recent Mineral Resource estimate, which is the subject of this audit, was prepared by Messrs. Robert Sim, P. Geo., and Bruce Davis, FAusIMM, in September 2009 (Sim and Davis, 2009), hereafter referred to as the "2009 resource model" or "2009 resource estimate". A probability shell was generated using Indicator Kriging (IK) to serve as a constraining wireframe model for the deposit. The estimate was made using a block model method, with grades interpolated using OK for Cu, Au, Ag, and Pd.

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In addition, potentially deleterious components were estimated into the blocks, including As, Hg, Sb, Bi, Cd, Fe, Ni, Pb, S, Se, Te, and Zn. These estimates were made using Inverse Distance Squared (ID2) weighting.

The Mineral Resource estimate, as reported by Sim and Davis (2009) is shown in Table 17-1.

The model was constructed using MineSite software, which is an off-the-shelf commercial package commonly used in the industry. Wireframe models of the topography and the DMZ were generated by New Gold personnel. The probability grade shell was created using IK, with some manual modifications afterward. The grade shell encloses blocks with a greater than 50% probability of grading greater than 0.25% Cu.

The database for the grade interpolations was entirely composed of diamond drill samples collected by New Gold over the period 2000 to the present.

TABLE 17-1 MINERAL RESOURCE SUMMARY

New Gold Corp. – New Afton Project

Cut-off Grade	Mtonnes	Cu%	Au g/t	Ag g/t	Pd g/t
(Cu%)	Mtonnes	Cu%	Au g/t	Ag g/t	Pd g/t
Measured
0.2	37.7	1.19	0.85	2.6	0.10
0.3	34.5	1.28	0.92	2.8	0.11
0.4	33.1	1.32	0.94	2.9	0.11
0.5	31.8	1.35	0.97	2.9	0.11
0.6	30.3	1.39	0.99	3.0	0.11
0.7	28.5	1.44	1.02	3.1	0.11
0.8	26.6	1.49	1.05	3.2	0.11
0.9	24.5	1.55	1.08	3.3	0.11
1.0	22.4	1.60	1.11	3.5	0.11
Indicated
0.2	40.4	0.69	0.52	1.8	0.08
0.3	27.6	0.90	0.68	2.3	0.10
0.4	23.0	1.01	0.76	2.6	0.11
0.5	20.0	1.10	0.82	2.8	0.11
0.6	17.7	1.17	0.87	2.9	0.11

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Cut-off Grade	Mtonnes	Cu%	Au g/t	Ag g/t	Pd g/t
0.7	15.4	1.25	0.92	3.1	0.11
0.8	13.4	1.32	0.96	3.2	0.11
0.9	11.5	1.40	1.00	3.4	0.11
1.0	9.8	1.48	1.04	3.5	0.12
Measured and Indicated
0.2	78.1	0.93	0.68	2.2	0.09
0.3	62.1	1.11	0.81	2.6	0.10
0.4	56.0	1.19	0.87	2.8	0.11
0.5	51.8	1.26	0.91	2.9	0.11
0.6	48.0	1.31	0.95	3.0	0.11
0.7	43.9	1.37	0.98	3.1	0.11
0.8	40.0	1.44	1.02	3.2	0.11
0.9	35.9	1.50	1.05	3.3	0.11
1.0	32.1	1.57	1.09	3.5	0.11
Inferred
0.2	38.9	0.51	0.38	1.3	0.07
0.3	21.7	0.73	0.52	1.8	0.08
0.4	16.0	0.87	0.62	2.1	0.09
0.5	12.8	0.98	0.70	2.4	0.09
0.6	10.3	1.09	0.78	2.6	0.10
0.7	8.5	1.19	0.85	2.8	0.10
0.8	7.3	1.26	0.90	3.0	0.10
0.9	6.2	1.33	0.95	3.2	0.10
1.0	5.2	1.41	0.99	3.4	0.10

Notes:

1.	CIM definitions were followed for Mineral Resources.

2.	Mineral Resources are estimated at a cut-off grade of 0.5% Cu.

3.	Mineral Resources are inclusive of Mineral Reserves.

MODEL GEOMETRY

A project survey grid has been established which is rotated 50° clockwise from the UTM grid, and the database, wireframes and block models are based on this grid. The block model consists of an array of 10 m x 10 m x 10 m blocks. Origin and model size parameters are shown in Table 17-2.

Note that the origin coordinates in Table 17-2 use the Gemcom convention. In Gemcom, the block model origin is the uppermost southwest corner (i.e., not the centroid) of the

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uppermost southwest block. Columns increment towards grid east, rows towards grid north, and levels downwards.

TABLE 17-2 BLOCK MODEL GEOMETRY

New Gold Inc. – New Afton Project

Origin:	X	3,000
	Y	1,650
	Z	5,750

Rotation:		0

Block Size:	X	10
	Y	10
	Z	10

Columns:		115
Rows:		55
Levels:		115

DATABASE

SAMPLES

The data for the wireframing and grade interpolations are entirely from diamond drill sampling. The database contains records for 276 holes, of which 28 are located outside the resource model area. The number of sampled intervals varies depending on the element. Key economic components such as Cu, Au, Ag, and Pd were assayed in virtually all samples, and numbered in the order of 47,000 to 49,000. Deleterious components, such as Hg, As, and Sb were analyzed in only 7,100 or so samples, while total S, Se, and Te in only approximately 2,600 samples.

A table of sample statistics used in the estimate is provided below in Table 17-3.

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TABLE 17-3 STATISTICAL SUMMARY OF SAMPLE ASSAY DATA
New Gold Inc. – New Afton Project

Element	No. of Samples	Total Length (m)	Min	Max	Mean	Std Dev
Copper (%)	48,670	101,588.9	0.01	13.70	0.325	0.695
Gold (g/t)	48,969	102,279.2	0.00	22.10	0.279	0.655
Silver (g/t)	46,977	98,134.9	0.00	1,880.00	0.971	8.896
Palladium (g/t)	48,971	102,283.2	0.00	14.10	0.048	0.141
Arsenic (ppm)	7,161	14,918.2	0.10	10,001.00	142.910	643.170
Iron (%)	7,153	14,902.2	0.62	21.78	5.010	1.909
Mercury (ppb)	7,064	14,724.4	2.00	94,600.00	635.020	2,882.070
Antimony (ppm)	7,161	14,918.2	0.00	2,365.00	23.830	94.970
Selenium (ppm)	2,651	5,296.3	0.10	36.40	2.620	2.980
Tellurium (ppm)	2,610	5,214.3	0.01	2.88	0.140	0.267
Bismuth (ppm)	7,159	14,914.2	0.01	190.00	4.986	10.776
Cadmium (ppm)	7,159	14,914.2	0.01	163.81	0.828	2.347
Nickel (ppm)	7,159	14,914.2	0.50	1,480.00	61.73	122.640
Lead (ppm)	7,159	14,914.2	0.01	410.00	22.561	23.655
Sulphur (%)	2,610	5,214.3	0.01	10.10	1.537	1.797
Zinc (ppm)	7,159	14,914.2	0.50	1,346.00	82.480	98.620

COMPOSITES

The sample data were composited to two-metre downhole lengths. The two-metre interval was selected because approximately 93% of the samples are two metres in length and because this length results in the least change to the original data. In Scott Wilson RPA’s opinion, this is a reasonable criterion for selection of the composite interval.

In reviewing the data, Sim and Davis (2009) noted that there were several sections of the drill holes that were not sampled (termed “missing intervals”). Some of the unsampled intervals were non-mineralized and deliberately ignored to save assay costs. Other missing intervals occurred in the database where holes were wedged. The upper portion of the pilot hole down to the wedge is carried in the record for that wedge to ensure that it plots at the appropriate depth. The samples for this interval are assigned to the pilot hole, and consequently, the upper portion of the wedge hole appears as

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“missing” in the sample record. Still other intervals are truly “missing” (i.e., they have been lost and are unaccounted for).

It is conventional practice to assign zero grade to missing intervals, however, there are several instances in the database where this would be unfairly conservative, and these should be ignored. Sorting out which missing intervals should be assigned zero grade is onerous and error-prone. Sim and Davis (2009) carried out a study to determine what effect ignoring all of the missing intervals had on the global resource estimate and found this effect to be negligible. Consequently, all missing intervals were ignored in the compositing. In Scott Wilson RPA’s opinion, this is a reasonable approach.

ESTIMATION DOMAINS

For the 2006 estimate, New Gold geologists had prepared wireframe models of what they perceived to be the principal mineralogical zones within the deposit. There were four principal zones of mineralization: supergene, mesogene, hypogene, and the C Zone. The hypogene zone was simply the primary sulphide mineralization, consisting of chalcopyrite and lesser bornite with accessory pyrite. The mesogene zone was defined as encompassing material that had begun to show signs of oxidation, typified by the emergence of chalcocite replacing primary Cu sulphides. The supergene zone was interpreted to be the region of the deposit where the Cu minerals had been entirely oxidized to native Cu or remnant chalcocite. The C Zone was (and still is) considered to be a separate hypogene sulphide body residing below the main body of mineralization.

A fifth domain, the DMZ, was recognized and partially modelled but was not as well understood as it is now. Further interpretive work by New Gold geologists has determined that this is a distinct post-mineral intrusive phase that can be identified in drill core, but had hitherto been considered an altered equivalent of the host units.

Additional wireframe modelling has also been completed on the fault zones which both bracket and transect the deposit. The two principal fault zones, the Hanging Wall and Footwall, have been modelled in some detail. Additional crossing structures and splays have been modelled to a somewhat lesser degree of completeness yet still serve to improve overall understanding of the controls to mineralization.

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For the updated model, a review was undertaken to determine if the various domains from the 2006 models could be used. The following conclusions were made:

·	The boundaries of the zones, as they appear in the drilling, are indistinct and somewhat subjective.

·	The geologists responsible for interpreting the zones had since left the company, and the wireframes had not been fully updated with the newer drill information.

·	There was scepticism and doubt as to the original interpreted zones and not enough time or resources to reinterpret them.

·	Statistical analyses and contact profiles demonstrated that there were no real differences in grade characteristics between the zones; therefore, constraining the grade interpolations on that basis did not appear to be relevant.

As a result, the decision was taken to construct a grade shell as the defining boundary of the deposit. The 2006 wireframes were retained for assignment of bulk densities to the blocks and for general characterization of the deposit.

The grade shell was derived from an IK model wherein material with a greater than 50% probability of grading higher than 0.25% Cu was defined as “in the zone”. From this, three principal estimation domains were established: the DMZ, the grade shell, and outside of either the grade shell or the DMZ.

Scott Wilson RPA inspected the probability shell and attempted to re-create it from the 2009 resource model composite data. A semi-variogram analysis was carried out using Sage, the same software used by Sim and Davis (2009). The variogram model derived from this analysis was built into a model in GEMS, and a new indicator model generated. Scott Wilson RPA notes that the GEMS software is more limited than MineSite and so it was not possible to exactly copy the search parameters. The resulting probability shell was compared to the 2009 resource model in plan and section views. For the most part, the two wireframes were similar in size and shape, however, there were some notable differences which have yet to be explained. As it happens, this turned out to be inconsequential, as a resource estimate created from the GEMS probability shell yielded almost identical results as those in the 2009 resource estimate. Scott Wilson RPA concludes that there is very little sensitivity in the model to small changes in the probability shell.

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The use of probability shells in resource estimation is an acceptable practice and, if prepared by experienced practitioners (as in this case), a reasonable approach. However, in Scott Wilson RPA’s opinion, it would be preferable to use deterministic models of the geological domains interpreted from lithological or mineralogical characteristics. In this case, time and cost constraints for completing the estimate left little choice. Reverting to the grade shell has not had a significant impact on either the Mineral Resource or Mineral Reserve estimates. The last three estimates for the New Afton deposits have been prepared using markedly different methods, which all have produced similar results. In Scott Wilson RPA’s opinion, this suggests that the estimate is quite robust and is based on sufficient drilling information to be relatively insensitive to changes in methodology.

In Scott Wilson RPA’s opinion, domains based on mineralogical parameters are useful for reasons other than the Mineral Resource estimate. Metallurgical recoveries are observed to vary between the different mineralogical zones. Similarly, the geotechnical characteristics of the rock mass are observed to have a measurable impact on the quality of the rock mass. Scott Wilson RPA recommends that for future block models these components be included in the modelling process, if possible.

CAPPING OF HIGH GRADES

Composite statistics for the economic components in the three principal domains are shown in Table 17-4. Histograms and probability plots are provided in Appendix 1.

Scott Wilson RPA notes that the distributions of all components are positively skewed and this generally results in overestimation of metal contents unless some remedial step is taken. Mean values generated from non-normally distributed data are often times biased in relation to the true average grade. A common practice in the industry is to cap grades at some predetermined level, and in so doing, limit the impact of the highest grade samples. The 2009 resource model incorporates grade caps that have been applied to the composited sampled data prior to running the grade interpolation. A summary of the caps applied is provided in Table 17-5.

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TABLE 17-4 TOP CUTS
New Gold Inc. – New Afton Project

Element	Top-Cut Limit			% Metal Loss in Model
Element	Inside Prob Shell	Outside Prob Shell	DMZ	% Metal Loss in Model
Copper	7%	3%	3%	-0.3
Gold	7 g/t	5 g/t	3 g/t	-1.2
Silver	40 g/t	15 g/t	10 g/t	-3.3
Palladium	2.5 g/t	2.0 g/t	0.6 g/t	-2.1
Arsenic	6,000 ppm			-1.5
Mercury	7,000 ppb			-1.7
Antimony	1,250 ppm			-2.4
Bismuth	80 ppm			-0.7
Cadmium	40 ppm			-5.3
Nickel	900 ppm			-0.2
Lead	200 ppm			-0.2
Iron	Not required			n/a
Zinc	1,000 ppm			-0.2

Scott Wilson RPA generally advises against placing caps on the composited samples in favour of capping the raw samples. However, in this case, because the compositing has had a minimal impact on the sample lengths, it is acceptable. Scott Wilson RPA reviewed the cap values and considers them to be within a reasonable range.

GEOSTATISTICS

Preparation of the 2009 resource model included a geostatistical analysis on the composited data in order to derive kriging and search parameters. Multi-directional correlograms for Cu, Au, Ag, and Pd were generated using Sage software. The data were partitioned into two domains: one inside and one outside the DMZ.

The variogram models derived from the analysis are shown in Table 17-5.

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TABLE 17-5 VARIOGRAM MODELS
New Gold Inc. – New Afton Project

				1st Structure			2nd Structure
Element	Nugget	S1	S2	Range (m)	AZ	Dip	Range (m)	AZ	Dip
Copper	0.2	0.26	0.54	52	203	78	491	82	45
	In/Out Prob Shell			46	22	12	272	97	-44
	In/Out Prob Shell			42	292	0	112	359	-7
Copper	0.3	0.547	0.153	50	339	64	538	348	64
	DMZ domain			37	104	16	91	315	-22
	DMZ domain			7	200	20	79	51	-13
Gold	0.3	0.29	0.41	39	66	38	475	79	57
	In/Out Prob Shell			27	327	12	237	103	-31
	In/Out Prob Shell			25	223	50	102	6	-11
Gold	0.4	0.35	0.25	26	275	-38	469	38	84
	DMZ domain			13	293	50	88	143	2
	DMZ domain			4	192	9	26	53	-6
Silver	0.4	0.226	0.374	72	62	46	569	75	44
	In/Out Prob Shell			31	200	36	277	80	-46
	In/Out Prob Shell			28	307	23	105	348	-3
Silver	0.4	0.336	0.264	33	236	-60	357	326	0
	DMZ domain			13	346	-11	302	56	70
	DMZ domain			5	262	28	29	56	-20
Palladium	0.45	0.341	0.209	48	13	61	609	85	53
	In/Out Prob Shell			24	239	21	292	104	-35
	In/Out Prob Shell			13	321	-19	117	8	-9
Palladium	0.399	0.474	0.127	48	116	13	190	102	70
	DMZ domain			16	338	73	163	22	-4
	DMZ domain			5	208	11	60	113	-19

Scott Wilson RPA conducted an independent geostatistical analysis using the same software (Sage) and the same composite database. The variogram models derived from this analysis were similar in many respects to those in Table 17-5, but with some significant differences. For the most part, the anisotropy directions and ranges were similar for the second structures of the models. The first structures, however, tended to differ substantially. In addition, the nugget effects were generally significantly lower than those derived for the 2009 resource model.

It is curious why the software (which, incidentally, derives the variogram models automatically without much user intervention) should generate different variogram

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models from, supposedly, the same data. However, it does not appear to be material to the global estimate. As mentioned above, Scott Wilson RPA ran grade interpolation models using an independently derived probability shell and variogram models and obtained results very similar to those obtained by Sim and Davis (2009). This suggests that the global estimate is insensitive to changes to the variogram models.

SEARCH PARAMETERS

For the economic elements in the DMZ domain, the search was a 300 m spherical ellipsoid, with a minimum of 8 and a maximum of 28 composites per estimate, and a maximum of seven composites allowed from any one drill hole. Search parameters were the same for the grade shell domain, except that for Au, and the estimates were limited to a minimum of six and a maximum of 15 composites, with a maximum of five allowed from any one drill hole. For outside of the probability shell, the composite constraints were identical to those for the DMZ, except that the search ellipsoid was reduced to 200 m x 200 m x 100 m. For all domains, the estimates were constrained to using only one hole per octant.

For the ID2 estimates (deleterious components), a 300 m spherical search was used, with the block estimates limited to a minimum of four and a maximum of nine composites, and again, a maximum of one hole per octant.

In Scott Wilson RPA’s opinion, the search parameters for the economic elements are reasonable and supported by the variography. The 300 m search used for the deleterious elements is somewhat aggressive. As a result, the estimates for these elements are likely to be lower in confidence level than the economic elements. This is particularly likely considering the lower number of samples for the deleterious elements.

BULK DENSITY

Average bulk density was assigned to the block model according to mineralization type. The average bulk densities were derived from 656 measurements collected from drill core over the course of exploration work on the Project. Table 17-6 shows how the assignments were made.

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TABLE 17-6 BULK DENSITY
New Gold Inc. – New Afton Project

Minzone Domain	Bulk Density (t/m3)
Supergene	2.58
Mesogene	2.59
Hypogene	2.62
Other (Py-Mt)	2.59

VALIDATION

Sim and Davis validated the 2009 resource block model using the following techniques:

·	Visual inspection and comparison of block and composite grades in sections and plans.

·	Hermitian Polynomial Change of Support analysis to check the degree of smoothing in the block grade interpolations.

·	Comparison with ID2 and Nearest Neighbour (NN) estimates.

·	Swath plots.

The validation exercises conducted by Sim and Davis (2009) indicated that the block grade interpolations were unbiased and honoured the local drill hole grades. In Scott Wilson RPA’s opinion, the validation work by Sim and Davis (2009) was done at a reasonable and appropriate level.

Scott Wilson RPA prepared an independent estimate for Cu, Au, Ag, and Pd and compared it to the 2009 resource estimate. As already stated in this report, a new probability shell was derived from an independently generated set of kriging parameters. This shell was then used, along with the DMZ solid, to assign domains to the blocks and composites. Geostatistical analyses were carried out to derive variogram models, which, as already mentioned, were somewhat different from those generated by Sim and Davis (2009). Capping, search parameters, and classification criteria (see subsection on Classification) were kept the same as for the original models.

The Scott Wilson RPA block model results at the 0.5% Cu cut-off are compared to the 2009 resource model in Table 17-7. There is very little difference between the two

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estimates, which, in Scott Wilson RPA’s opinion, provides additional validation of the 2009 resource model. Scott Wilson RPA also notes that the global estimate appears to be quite insensitive to variations in either the probability shell or the variogram models.

CLASSIFICATION

The estimate has been classified according to the CIM Definition Standards for Mineral Resources and Mineral Reserves as adopted by the CIM Council on December 11, 2005 (CIM definitions). Criteria for classification were as follows:

MEASURED

Assigned to blocks with Cu grades estimated by a minimum of three drill holes with composites at an average spacing of 30 m (nominal drill spacing of approximately 50 m).

INDICATED

Assigned to blocks with Cu grades estimated by a minimum of three drill holes with composites at an average spacing of 50 m (nominal drill spacing of approximately 80 m).

INFERRED

Assigned to blocks that do not meet the criteria for Measured or Indicated but are within 50 m of a drill hole.

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TABLE 17-7 MINERAL RESOURCE ESTIMATE COMPARISON
New Gold Inc. – New Afton Project

Scott Wilson RPA Model					2009 Mineral Resource Model					Percent Difference (%)
Tonnage (Mt)	Cu (%)	Au (g/t)	Ag (g/t)	Pd (g/t)	Tonnage (Mt)	Cu (%)	Au (g/t)	Ag (g/t)	Pd (g/t)	Tonnage	Cu	Au	Ag	Pd
Measured
31.8	1.35	0.97	2.90	0.11	32.7	1.34	0.97	2.78	0.11	-2.8	0.7	0.0	4.3	0.0
Indicated
20.0	1.10	0.82	2.80	0.11	19.0	1.11	0.87	2.72	0.11	5.3	-0.9	-5.7	2.9	0.0
Measured + Indicated
51.8	1.26	0.91	2.90	0.11	51.7	1.26	0.93	2.76	0.11	0.2	0.0	-2.2	5.1	0.0
Inferred
12.8	0.98	0.70	2.40	0.09	12.4	0.99	0.74	2.38	0.09	3.2	-1.0	-5.4	0.8	0.0

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CUT-OFF GRADE

Sim and Davis (2009) used a 0.25% Cu cut-off for the probability shell, and applied a 0.5% Cu cut-off to the block grades for reporting purposes. The cut-off used for the Feasibility Study resource estimate was US$10/t (Hatch, 2007). For the present model, the cut-off is based on Cu grade only (0.5% Cu), which will likely result in a reduction in the overall size of the estimate.

In Scott Wilson RPA’s opinion, there is significant value in the other metals, particularly Au, which is not being given full credit by using a Cu-only cut-off. New Gold’s Life of Mine model shows that contribution to revenue is as follows; Cu = 70%, Au = 29% and Ag = 1%. On the recommendation of Scott Wilson RPA, Sim and Davis (2009) prepared a report of the Mineral Resources using a Cu-equivalent (CuEq) cut-off. Copper equivalence was calculated based on the following equation:

CuEq = Cu_% + (DolVal_Au / DolVal_Cu) * Au_g/t

Where:

CuEq = copper equivalent grade

Cu_% = copper grade in percent

Au_g/t = gold grade in g/t

DolVal_Cu = $/% Cu = $55.12 (based on $US2.50/lb Cu)

DolVal_Au = $/g Au = $28.94 (based on $US900/oz Au)

The revised Mineral Resources are summarized in Table 17-8.

TABLE 17-8 MINERAL RESOURCE SUMMARY (CUEQ CUT-OFF)

New Gold Corp. – New Afton Project

Cut-Off	Tonnage	CuEq	Cu	Au	Ag	Pd
(% CuEq)	(Mt)	(%)	(%)	(g/t)	(g/t)	(g/t)
Measured
0.2	43.1	1.46	1.06	0.77	2.3	0.10
0.3	37.3	1.66	1.20	0.87	2.6	0.11
0.4	35.1	1.74	1.26	0.91	2.8	0.11
0.5	33.7	1.79	1.30	0.94	2.8	0.11
0.6	32.6	1.83	1.33	0.96	2.9	0.11
0.7	31.6	1.87	1.36	0.98	2.9	0.11
0.8	30.5	1.91	1.38	1.00	3.0	0.11
0.9	29.5	1.95	1.41	1.02	3.1	0.11
1.0	28.2	2.00	1.45	1.05	3.1	0.12

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Cut-Off	Tonnage	CuEq	Cu	Au	Ag	Pd
(% CuEq)	(Mt)	(%)	(%)	(g/t)	(g/t)	(g/t)
Indicated
0.2	63.8	0.70	0.49	0.41	1.3	0.07
0.3	39.4	0.99	0.69	0.56	1.9	0.09
0.4	29.9	1.20	0.85	0.67	2.2	0.10
0.5	25.2	1.34	0.95	0.74	2.5	0.10
0.6	22.1	1.45	1.03	0.80	2.7	0.11
0.7	19.9	1.54	1.10	0.84	2.8	0.11
0.8	18.1	1.62	1.15	0.89	2.9	0.12
0.9	16.6	1.69	1.20	0.93	3.0	0.12
1.0	15.2	1.76	1.25	0.97	3.1	0.12
Meas. & Ind.
0.2	107.0	1.01	0.72	0.55	1.7	0.08
0.3	76.7	1.31	0.94	0.71	2.2	0.10
0.4	65.0	1.49	1.07	0.80	2.5	0.10
0.5	58.8	1.60	1.15	0.85	2.7	0.11
0.6	54.7	1.68	1.21	0.89	2.8	0.11
0.7	51.5	1.74	1.26	0.93	2.9	0.11
0.8	48.7	1.80	1.30	0.96	3.0	0.12
0.9	46.1	1.85	1.34	0.99	3.0	0.12
1.0	43.3	1.91	1.38	1.02	3.1	0.12
Inferred
0.2	72.2	0.50	0.34	0.31	0.9	0.06
0.3	38.4	0.73	0.51	0.43	1.3	0.07
0.4	25.2	0.94	0.66	0.54	1.7	0.08
0.5	19.0	1.10	0.78	0.61	2.0	0.09
0.6	15.1	1.25	0.89	0.68	2.2	0.09
0.7	12.6	1.37	0.98	0.74	2.4	0.10
0.8	10.7	1.48	1.07	0.80	2.6	0.10
0.9	9.2	1.59	1.14	0.85	2.8	0.10
1.0	8.2	1.66	1.20	0.89	2.9	0.11

Table 17-8 demonstrates that the CuEq cut-off adds a modest but measureable amount to the Mineral Resource estimate (see Table 17-1). The total Measured and Indicated tonnage, at the 0.5% CuEq cut-off, increased by approximately 13%, with a partially offsetting drop in grades. The overall metal content for Cu and Au at that cut-off increase by 4% and 6%, respectively.

The 0.5 % Cu (or CuEq) cut-off, at US$2.50/lb Cu, represents a gross contained metal value of US$27.55/t. Using the parameters in the New Afton financial model, Scott Wilson RPA estimates that for ore grades of 0.5% Cu, 0.4 g/t Au, and 1.0 g/t Ag (i.e.,

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precious metal grades that are roughly proportional to the average Cu grade in the deposit), the NSR value of the rock is C$33.74/t. The estimated cash costs for LOM at New Afton are C$17.19/t, suggesting that the 0.5% Cu cut-off is very conservative, particularly for a Mineral Resource estimate. At a 0.3% Cu grade, the same NSR analysis indicates that the NSR value is C$20.24, which compares more favourably to the LOM cash costs per tonne.

In Scott Wilson RPA’s opinion, a cut-off in the order of 0.3% Cu is probably more appropriate for the Mineral Resource estimate than 0.5% Cu. At the 0.3% Cu-only cut-off, the additional metal values contributed by Au and to a much lesser degree, Ag, are in the order of an additional 0.1% CuEq. This implies that the CuEq cut-off would be in the order of 0.4% CuEq. Comparison between Tables 17-1 and 17-8 indicates that the tonnage reporting at the 0.3% Cu cut-off is roughly equal to that at the 0.4% CuEq. Consequently, Scott Wilson RPA recommends that the Mineral Resources be quoted at a CuEq cut-off of 0.4% CuEq.

COMPARISON WITH THE PREVIOUS ESTIMATE

Table 17-9 compares the 2009 Mineral Resource estimate, at the 0.4% CuEq cut-off, with the 2006 Mineral Resource estimate, at the $10/t cut-off, as reported in the Feasibility Study (Hatch, 2007). The tonnage of the Measured Resources dropped, while there was a significant increase for the Indicated Resources. Total tonnage of the Measured and Indicated Mineral Resources did not change appreciably, nor did the total metal contents. The Inferred Resources essentially tripled in volume, with a marked decrease in grades.

In Scott Wilson RPA’s opinion, the variance between the Measured plus Indicated Mineral Resources in the two estimates is not significant, and can be explained by differences in methodology. The changes in the Inferred Mineral Resources are due to additional drilling has expanded the C Zone, resulting in an increase in total Inferred tonnage. Additional drilling has also likely affected the resources in the area of the Hanging Wall Zones.

In Scott Wilson RPA’s opinion, the difference between the 2006 and present estimates is not significant on a global basis and should not make a material difference to the Mineral Reserve estimate.

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TABLE 17-9 2009 VS. 2006 ESTIMATE COMPARISON

New Gold Corp. – New Afton Project

2006 Estimate ($10/t C/O)					2009 Estimate (0.4% CuEq C/O)					Pct Diff					Pct Metal Diff
Mt	Cu %	Au g/t	Ag g/t	Pd g/t	Mt	Cu %	Au g/t	Ag g/t	Pd g/t	Mt	Cu	Au	Ag	Pd	Cu	Au	Ag	Pd
Measured
43.2	1.12	0.83	2.68	0.10	35.1	1.26	0.91	2.75	0.11	-18.8%	12.5%	9.6%	2.6%	8.0%	-8.6%	-10.9%	-16.6%	-12.3%
Indicated
22.4	0.84	0.66	2.42	0.09	29.9	0.85	0.67	2.24	0.10	33.5%	0.7%	1.1%	-7.5%	6.7%	34.4%	34.9%	23.4%	42.4%
Measured + Indicated
65.6	1.02	0.77	2.59	0.10	65.0	1.07	0.80	2.51	0.10	-0.9%	4.4%	3.4%	-3.0%	6.1%	3.5%	2.5%	-3.8%	5.1%
Inferred
7.9	0.96	0.88	1.55	n/a	25.2	0.66	0.54	1.69	0.08	219.0%	-31.6%	-38.8%	9.0%	n/a	118.3%	95.4%	247.8%	n/a

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MINERAL RESERVES

For the 2007 Feasibility Study, AMC Consultants Pty Ltd. (AMC) prepared a Mineral Reserve estimate based on the block caving method. The key inputs and processes used to develop and validate the estimate are described below. The 2007 Mineral Reserve estimate was based upon a geological block model prepared by Scott Wilson RPA which formed the basis of the Mineral Resource estimate published by New Gold in September 2006.

The 2007 reserve estimation was completed based upon the use of two separate modelling approaches (Cave-Sim and Gemcom PCBC). In September 2009, New Gold provided a new resource model which was reviewed and accepted by Scott Wilson RPA as described above. AMC again used the PCBC modelling software to generate a Mineral Reserve estimate. Based upon the similar results from the two software packages in the 2007 estimate, this 2009 estimate was based solely on the PCBC model for its ease of use.

Scott Wilson RPA has reviewed the assumptions and results of the estimation process and concurs that the estimate is consistent with the CIM definitions and NI 43-101. The Mineral Reserve estimate is summarized in Table 17-10 and described in more detail in the following subsections.

TABLE 17-10 MINERAL RESERVE ESTIMATE – SEPTEMBER 2009

New Gold Inc. – New Afton Project

Category	Tonnage (M tonnes)	Cu (%)	Au (g/t)	Ag (g/t)
Probable	47.4	0.95	0.69	2.03

Notes:

1.	CIM definitions were followed for Mineral Reserves.

2.	Mineral Reserves are estimated at an NSR cut-off grade of $19/t.

3.	Mineral Reserves are estimated at US$2.00/lb Cu and US$800/oz Au and a C$/US$ exchange rate of 0.85.

4.	Metallurgical recoveries for copper vary from 69% to 89%.

2007 MINERAL RESERVE ESTIMATE

The 2009 Mineral Reserve estimate is based on the fundamental assumptions and parameters that were developed for the 2007 Mineral Reserve estimate and are

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described below. In preparing the 2007 estimate, AMC used geotechnical data and geological interpretations of faults and other structures provided by New Gold.

A detailed analysis of the deposit was first carried out to identify the most suitable positions of the block cave extraction levels and the extent of the area to be undercut to induce caving (the block cave footprint). During this analysis a large number of iterations were carried out.

After selection of appropriate extraction levels and footprints and an overall development sequence, taking into account geotechnical and practical mining constraints, a detailed production schedule was developed using commercially available block cave modelling software (PCBC).

An alternative software program (Cave-Sim) was also used to model material flow in the cave and produce the production estimates for the Feasibility Study. Sensitivities to changes in key parameters influencing the reserve estimate were modelled using both PCBC and Cave-Sim.

INITIAL REVIEW OF RESOURCE GRADE DISTRIBUTION

Those parts of the deposit suitable for mining were initially identified by visual inspection of the grade distribution within the deposit, and by analysis of the value generated by mining various areas. The Hanging Wall Zones are isolated from the Main Zone and are low grade. Because of their location, they cannot utilize the underground infrastructure required for mining the Main Zone and would require separate access, ore handling and ventilation systems. These zones were considered to be uneconomic.

The Main Zone was split into east and west blocks at 3525E, which coincides with the approximate position of a low-grade zone running through the deposit. Both the east and west portions of the Main Zone were used for further analysis to determine the most suitable extraction levels and footprints.

SELECTION OF EXTRACTION LEVEL AND FOOTPRINT

AMC reviewed a number of possible undercut positions by visual inspection of the resource. The block size in the resource model was reset to 15 m x 15 m x 20 m high to match a preliminary estimate of drawpoint spacing. A series of block models were then

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created, one for each potential undercut position. In each model, blocks below the undercut were removed from the model, while those above the undercut were subjected to a mixing algorithm to simulate the effect of drawing material from drawpoints.

AMC added a value field (Net Smelter Return, or NSR, Value) to the diluted model. The NSR Value is an estimate of the theoretical value received for the concentrate produced from one tonne of resource, net of all costs and losses incurred in the off-site transportation and processing of concentrate.

After calculating the unit NSR Value of the material in each diluted block, an amount equivalent to $14 per tonne was deducted from the NSR Value to reflect the mining, onsite processing, and administration costs. No account was taken of the time value of money. Columns of blocks having a positive cumulative NSR Value were assembled, progressing up from the undercut. The final height of each column was determined by the position of the last block to add value to the column. An amount reflecting a preliminary estimate of the cost of establishing the portions of the undercut, extraction level, drawpoint, and drawbell necessary to mine each column was subtracted from the total column value.

A perimeter or footprint was drawn on the undercut level around groups of profitable columns. The footprint was adjusted to ensure that a minimum north-south span was maintained. A number of footprints were prepared for each model to test the sensitivity of the Project to minimum span criteria of 105 m, 90 m, and 75 m. Unprofitable columns were added to the footprint where necessary, to achieve the minimum span. The height of unprofitable columns was limited to 60 m to minimize the mining of unprofitable material, while ensuring full cave establishment.

A large number of possible extraction level options were initially tested, leading to the conclusion that the most suitable arrangement was to establish three panels – Block 1 (B1), Block 2 (B2), and Block 3 (B3) - with B1 and B2 on the same elevation and B3 on a lower elevation. B3 would mine the down-plunge western extension of the deposit and could mine underneath B2 if necessary.

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Further work was then carried out to establish the optimum elevation of the B1 and B2 undercut, the western limit of B2, and the elevation and eastern limit of the B3 undercut. The undercut level for B1 and B2 was set at a nominal elevation of 5,085 m and for B3, at 4,965 m.

To enable a single cut-off parameter to be used, AMC added an NSR Value field to the Combined Resource Model. Each of the different rock types - supergene, mesogene, and hypogene - have different recoveries and concentrate grades, which impact on the NSR Value calculation.

PRODUCTION ESTIMATES USING PCBC

An assessment of the recoverable tonnage and grade from the deposit was carried out on behalf of AMC by Stratavision Pty Ltd. (Stratavision). The PCBC software program was used for this assessment. Key inputs to the PCBC modelling process include the estimated shape of the draw cones, the percentage of fines, the mixing horizon and the sequence of extraction. Three cases were modelled using a set of Expected, Best Credible, and Worst Credible parameters.

PCBC PRODUCTION SEQUENCING

An extraction sequence starting in the southwest corner of B2 and progressing to the northwest corner of B1 was used in all cases. B3 was scheduled to commence after completion of the western drawpoints in B2. The undercut face was maintained at a nominal angle of 45° to the strike of the orebody. The production schedule for B3 was run independently of B1 and B2. In reality, the far eastern drawpoints of B3 will incur slightly more dilution than indicated, due to the presence of caved waste in the depleted western portion of B2.

The scheduling method used takes the total required tonnage and divides it amongst all the active drawpoints in proportion to the amount of material remaining in each draw column. Thus the higher columns are drawn down faster and the lower ones more slowly. The aim is to close off all the drawpoints at approximately the same time. This is more difficult to achieve if there are significant variations in economic column heights between drawpoints. This was experienced in B1 where the far eastern drawpoints had to be “trimmed” to a maximum Height of Draw (HOD) of 350 m

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PRODUCTION ESTIMATES USING CAVE-SIM

In addition to using PCBC, an assessment of the recoverable tonnage and grade from the deposit was carried out using the three-dimensional particle flow code Cave-Sim. The code is broadly based on the concept of cellular automata, in which cells are considered as discrete objects that are categorized by individual "state" parameters that evolve dynamically. This allows very large systems to be studied with very simple discrete equations that are calibrated against empirical knowledge. The three dimensional cellular automata in Cave-Sim consider particle friction, particle size distributions, and state of stress. The user has full control over draw rates from individual drawpoints and tracking of the instantaneous and cumulative grade and particle size distribution through each drawpoint.

COMPARISONS BETWEEN PCBC AND CAVE-SIM RESULTS

In the Expected and Best Credible cases, Cave-Sim modelling indicates marginally higher levels of dilution than PCBC resulting in higher tonnages (+5%) and lower grades (-2% to -3%). However, the differences are small and within the accuracy of the estimating process.

ADJUSTMENT FOR ORE LOSSES

Both PCBC and Cave-Sim programs provide estimates of the tonnage and grade of the material that can potentially be recovered under a given set of draw control rules. In practice, events will occur that limit the ability to achieve these estimates. Adjustments to the schedules resulted in an overall reduction of approximately 4% in the tonnage output.

2009 MINERAL RESERVE ESTIMATE

A Mineral Reserve estimate was compiled by AMC based upon a revised resource model as described in the Mineral Resource section of this report. The 2009 Mineral Reserve estimate was based upon the use of the PCBC software for estimation of the mineral reserves within the planned block cave operation. The undercut elevations were unchanged from the 2007 Mineral Reserve estimate, however, the operating cost assumptions for the drawpoint shut-off grade were revised to match the current cost estimates. PCBC parameters were the same as the “expected” case in the feasibility study work.

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A lower grade pillar previously left between the B1 and B2 blocks has been mined and the B2 block has been expanded by rows of drawpoints to the west. This expansion to the west reduces the amount of material in the B3 block bring some production ahead and reducing haulage costs. There are two rows of B3 drawpoints which are located beneath the B2 cave.

The PCBC program includes dilution material and the draw height has been limited to a maximum of 350 m to account for the losses that will occur due to less than ideal draw control, damage to drawpoints and the loss of drawpoints that have not reached the shut-off grade at the end of the mine life.

The drawpoint production predicted from the PCBC model is shown in Table 17-11 along with the resource classification of material within the total.

TABLE 17-11 DRAWPOINT PRODUCTION BY RESOURCE CLASSIFICATION

New Gold Inc. – New Afton Project

	Mt	Cu %	Au g/t	Ag g/t	As ppm	Hg ppb	NSR C$
Measured	33.5	1.03	0.74	2.21	306	903	52.66
Indicated	12.6	0.77	0.58	1.66	265	970	39.82
Subtotal	46.1	0.96	0.70	2.06	295	921	49.16
Inferred	0.8	0.79	0.52	1.44	189	961	37.17
Unclassified	0.1	1.04	0.54	1.73	200	1,533	42.68
Subtotal	1.0	0.82	0.52	1.48	191	1,041	37.94
Total	47.0	0.96	0.69	2.05	293	923	48.92

In addition to the material generated from the drawpoints, there will be resources extracted in the course of mine development. This material and the resource classification is shown in Table 17-12.

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Table 17-12 Tonnage and Grade of Material from Mine Construction Activities

New Gold Inc. – New Afton Project

	Mt	Cu %	Au g/t	Ag g/t	As ppm	Hg ppb	NSR C$
Measured	0.187	0.87	0.78	1.66	85	388	49.45
Indicated	0.170	0.60	0.61	1.68	130	612	35.69
Subtotal	0.357	0.74	0.70	1.67	106	495	42.90
Inferred	0.004	0.20	0.28	0.66	275	848	13.22
Unclassified	-	-	-	-	-	-	-
Subtotal	0.004	0.20	0.28	0.66	275	848	13.22
Totals	0.361	0.74	0.70	1.66	108	499	42.55

There will be inferred and unclassified mineral resource material which will be mined from the block cave, and segregation of the material is not possible. The inferred and unclassified material is considered to carry some grade, though the material has insufficient certainty to be Measured or Indicated Resource. For the purposes of estimating the Mineral Reserves, which by CIM definitions include diluting materials, Scott Wilson RPA has included the tonnage of this inferred and unclassified material with the grades reduced by 50%. Scott Wilson RPA considers this material to be mineralized dilution, which will be included in the Mineral Reserve estimate and within the production plan.

CLASSIFICATION OF MINERAL RESERVES

There is a large volume of the material within the planned cave that has a Measured Resource classification. However, due to the uncertainty associated with estimating material movement within the cave, and the absence of any historical production versus forecast reconciliation, a Probable classification has been applied to the reserve because of the uncertainty.

MINERAL RESERVE ESTIMATE

The Mineral Reserve estimate for the New Afton Project is presented in Table 17-13. The Mineral Reserve has been estimated using the resource block model prepared by New Gold and confirmed by Scott Wilson RPA.

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Table 17-13 Mineral Reserve Estimate

New Gold Inc. – New Afton Project

	Mt	Cu %	Au g/t	Ag g/t	As ppm	Hg ppb
PCBC Resources	46.1	0.96	0.70	2.06	295	921
Internal dilution	1.0	0.41	0.26	0.74	96	521
Development	0.36	0.74	0.70	1.67	106	495
Total	47.4	0.95	0.69	2.03	290	911

Scott Wilson RPA is of the opinion the Mineral Reserve estimate conforms to the CIM definitions.

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18 OTHER RELEVANT DATA AND INFORMATION

Scott Wilson RPA is not aware of any other relevant data or information that should be included in this Technical Report.

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19 ADDITIONAL REQUIREMENTS

The New Afton deposit is being developed as a block cave mining operation to produce 4 Mtpa of copper-gold ore for processing in a flotation plant for the recovery of copper in concentrates and gold in concentrates. Underground development is underway using New Gold crews which replaced the mining contractor crews and resumed development in February 2009. There is supporting infrastructure at the site including access, offices, staff, equipment, and a concrete batch plant. The new mill building has been erected and the grinding mills are mechanically installed. No significant surface construction was underway at the time of the site visit and the focus is on underground development and project planning and engineering. This report considers the development of the Project from the end of the second quarter of 2009 onwards.

Previous underground mining studies related to the New Afton Project were completed by AMC as part of the Feasibility Study prepared by Hatch Ltd. (Hatch) in 2007. AMC completed additional engineering work subsequent to the Feasibility Study. AMC based its work on geotechnical, hydrological and geological information, including the Mineral Resource estimates, provided either directly by New Gold or other consultants involved in the Feasibility Study.

In the preparation of this Technical Report, Scott Wilson RPA has reviewed the mine design work by AMC, as well as the capital and operating cost estimates prepared for and after the 2007 Feasibility Study. It is Scott Wilson RPA’s opinion that the precision and level of detail applied to the design, scheduling, and cost estimation work is appropriate and consistent with the level of engineering and the state of the Project development. Scott Wilson RPA has used the design and descriptions as generated by AMC as the basis for this section of the report.

In the Feasibility Study, a staged start-up over a period of years was proposed, however, with the changes to the Project execution plan, the strategy is to move directly to the planned production rate of 4 Mtpa.

Development of the underground mine is the longest lead time item in the Project development. The Project construction schedule has been designed such that

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interconnections between independent development areas are established as soon as possible. Refuge chambers will be installed at suitable locations in each development area until the interconnections are established.

Initial ore from development, undercutting, and production drawpoints will be stockpiled in a temporary run-of-mine stockpile area adjacent to the mill. Milling operations were scheduled to commence in month 37, when the mine production and surface stockpile will provide sufficient ore for the mill to ramp up to 4 Mtpa over a three-month period.

MINING OPERATIONS

MINE DESIGN

During the mining method selection process for the feasibility study, studies were carried out to determine the most suitable production rate the Project. Initially, a staged production build was proposed starting with an annualized rate of 1.6 Mtpa for a nominal two-year period and then increasing to a maximum production rate of 4.0 Mtpa. With the break in the construction schedule caused by New Gold’s response to the global financial crisis, there has been time to advance the planning and mine development and the current strategy is to commence operations and ramp up to the 4.0 Mtpa rate immediately after start-up.

MINING METHOD

Block caving is proposed for the New Afton deposit after consideration of the following mining methods:

·	Non-Caving Methods

·	Open pit mining

·	Sublevel open stoping (SLOS) with cemented backfill

·	SLOS with uncemented backfill

·	Post pillar cut & fill

·	Caving Methods

·	Core & shell (open stoping with mass blasting of pillars)

·	Sublevel caving (SLC)

·	Block caving

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AMC concluded that open pit mining was uneconomical, open stoping methods would not be practical, and caving methods that started from the top (sublevel caving) would commence in the poorest ground conditions and in the supergene ore. Therefore, block caving was chosen because this method starts from the bottom and is conducive to large-scale low-cost mining.

A number of block cave footprints were prepared to test the sensitivity of the Project to minimum span criteria and a large number of possible extraction level options were initially tested. It was concluded that the most suitable arrangement was to establish three panels, with Block 1 (B1) and Block 2 (B2) on the same elevation and Block 3 (B3) on a lower elevation. The locations of the blocks relative to the open pit are shown in Figure 19-2.

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GEOMECHANICS

Geotechnical investigations of the New Afton deposit were completed by AMC for the Feasibility Study. AMC compiled a report setting out the relevant geotechnical aspects of the Project. The key geotechnical considerations relating to the design and operation of the Project are summarized below.

GEOTECHNICAL DOMAINS

The deposit is divided into three geological domains, based on the nature of the copper mineralization:

·	Hypogene (primary): Chalcopyrite ± chalcocite, with silicification and K-spar flooding, and variable calcite veins.

·	Mesogene (transition): Chalcopyrite, chalcocite, bornite, plus calcite veins and argillic alteration.

·	Supergene: Native copper, chalcocite, plus variable hematite and argillic alteration.

The following geotechnical domains were defined:

·	Supergene mineralization

·	Transitional (mesogene) mineralization

·	Primary (hypogene) mineralization

·	Fault Impacted Zone

·	Footwall (waste) rock.

With further drilling and underground exposures, the geotechnical domains may be refined in future.

The Fault Impacted Zone (FIZ) is a narrow (up to 20 m wide) zone of extremely poor ground conditions immediately adjacent and parallel to the Hanging Wall Fault. In addition to geotechnical domains, there is a picrite body to the south of the Hanging Wall Fault in close proximity to the orebody for a short strike length. Rockmass properties for this material were determined separately.

Alteration is highly variable and complex and is an important feature influencing rock mass quality. Alteration has not been explicitly modelled.

The structural geology of the deposit is complex. A number of different faulting orientations and styles are recognized (and have been wireframed). Lamprophyre and

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diabase dykes also intersect the deposit. Tertiary volcanic and sedimentary rocks lie to the north of the deposit, but do not impact directly on the underground project, although these areas will be partially impacted by surface settlement.

The most prominent fault sets which impact on mine design are:

·	Hanging Wall Fault Zone

·	Footwall Fault

·	Northeast faults

·	Northwest faults

·	North faults

·	Various picrite intrusives associated with major fault zones

Microfracturing is pervasive throughout the deposit and is most intensely developed where alteration is most intense. The intensity of microfracturing has not been recorded in geological logs. However, it is expected that microfracturing will prove to be a significant factor in cave development and fragmentation, and to a lesser degree in development stability.

Rock quality varies broadly from extremely poor to good, with an average of fair. The best ground conditions are found within the hypogene mineralization and the worst, within the intensely altered FIZ, the Hanging Wall Fault Zone, the adjacent picrite, and where major faults intersect the supergene mineralization.

GEOTECHNICAL DATABASE AND MODEL

There is a geotechnical database which includes the majority of the total drilling. Drilling coverage is good within and close to the mineralized zones, but decreases as the distance from the main mineralized areas increases. Specific drilling has provided geotechnical information in areas of key mine infrastructure. There is clear evidence of deterioration of the core over time, particularly in the weaker materials.

All geotechnical logging by New Gold has been based upon Bieniawski’s original 1976 version of his rock mass rating system, the RMR76. AMC reviewed the raw geotechnical data provided by New Gold, and generated RMR76 values during post-processing. The resulting database has been incorporated into a geotechnical block model.

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The New Afton drill hole database includes the following geotechnical data:

• Core recovery and RQD

• Fractures per metre

• Jointing intensity and condition (as per Bieniawski, 1976)

• Joint roughness (recorded separately to joint condition)

• Fracture / faulting code with core angle

• RMR76 (Bieniawski) – calculated field from the above

• RMR90 (Laubscher) – estimated from RMR76.

A database of small-scale structures (joints, fractures and veins) has been prepared from surface pit and underground mapping, and, based on a limited program of laboratory strength testing and field point load index testing, the rock mass unconfined compressive strengths were estimated. The virgin rock stress was measured during the study by overcoring of Hollow Inclusion (HI) cells in hypogene mineralization, exposed in the underground exploration crosscut.

To facilitate an empirical assessment of cavability, a conversion to Laubscher’s RMR (1990 version) was completed based on relogging of selected boreholes. This correlation is considered satisfactory for broad indications of behaviour, such as in a cavability assessment.

A geotechnical block model was prepared for both the mineralized zones and the host rocks. The block model is based upon simple kriging, with search ellipses based on geometry and statistical analyses, and criteria applied for minimum samples per cell. The model contains blank cells where there is insufficient drilling to satisfy the search criteria. Underground exposures provided a good opportunity to compare the block model with actual conditions, and the comparisons were good on the scale of groups of cells (tens of metres).

CAVABILITY

The cavability of a deposit defines the ability of the orebody and overlying rock mass to cave freely and spontaneously, once undercut to a sufficient dimension. Experience has shown there is a relationship between the competence of a rock mass and the dimensions of the undercut required to initiate and sustain caving. One of the design challenges is to determine the undercut dimensions required for caving of the given rock mass and to relate that to the orebody geometry available.

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Once initiated, caving is ideally a continuous process, providing sufficient void remains below the cave back to enable new failures to occur. This void is created by continuous production from the base of the cave, drawing down the broken caved material.

At New Afton, the dominant failure mechanism is projected to be structural collapse of the rock mass, i.e., there are sufficient natural planes of weakness in a variety of orientations for the rock mass to be unstable when undercut by a sufficient area.

The cavability of the New Afton deposit is represented by two end-members of a range of changing orebody geometries and ground conditions. In the east, the deposit is relatively wide, close to surface, and with a higher proportion of poorer ground conditions. Outside the FIZ domain, some of the poorest ground conditions are located in the Supergene ore at the far eastern ends of the deposit. Conversely, in the west, the hypogene orebody is more narrow, deeper, and located within some of the most competent ground conditions in the deposit.

As mentioned earlier in this section, it is proposed that the deposit will be mined using three block cave panels, B1, B2, and B3. Figure 19-3 shows the undercut footprints (blue outlines) for B1 and B2. Using these outlines the hydraulic radii (area/perimeter) for the three panels are as follows:

• B1 and B2: HR = 34 m

• B3: HR = 31 m

Based upon experience, there is no doubt the eastern areas of the orebody will cave readily. Further assessment of cavability is, however, required for the central and western areas of the deposit, especially where the orebody geometry narrows in the far west (and at depth). The cavability of the New Afton deposit has been assessed using two techniques: empirical and numerical stress analysis.

The empirical assessments of cavability indicate adequate cave development in B1 and B2, and marginal cavability for B3. Although B3 is indicated to be cavable, from the Laubscher Cavability Chart (Figure 19-4), the margin from the indicated “non-caving” zone is small. The north-south span for B3 in particular is recognized as a potential issue.

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The Abaqus non-linear finite element stress analyses indicated that a minimum north-south span of 95 m was required and that caving would initiate at a hydraulic radius (HR) of ≈ 32 m. This was based on hypogene rock properties. Again, this provides for adequate cavability for B1 and B2, but marginal cavability for B3.

The role of major structures has not been considered in either the empirical or numerical assessments of cavability. Major faults will assist cave development, especially in B1 which contains a high proportion of the weaker mesogene rock mass. The structural environment associated with B3 is not currently well defined and it is recommended that this be a focus for future data collection and investigative programs. B3 production is significantly later than B1 and B2, providing time for such studies.

UNDERCUTTING SEQUENCE

A number of undercutting sequencing alternatives were evaluated, but the sequence selected was B2 first, followed closely by B1, both adopting a southwest to northeast undercut direction. B3 (far west, and deeper) would be mined last.

The preferred undercutting sequence included consideration of the predominant structural fabric, the principal stress direction and the results of numerical modelling. The numerical modelling indicated that a significant zone of yield is likely to occur beneath the undercut level, as caving extends upwards. Damage zones also extend a significant distance into the surrounding rock mass adjacent to the caves, and suggest the undercuts should extend 30 m to 40 m ahead of drawbell construction. This has been implemented in the Project schedules.

CAVE CONTINGENCY PLANNING

Two options are considered to mitigate any potential risks of poor cave performance:

• Pre-conditioning, and/or

• Cave assistance.

Pre-conditioning refers to the use of various techniques to generate additional fracturing in the rock mass, to reduce rock mass strength, prior to the commencement of undercutting. Cave assistance refers to methods used to weaken the rock mass within the desired cave column, after caving has commenced, in response to areas of poor cave performance. It is not envisaged that pre-conditioning will be required in B1

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and B2 at New Afton. However, pre-conditioning using blasting or hydraulic fracturing techniques is an option for B3.

FRAGMENTATION

An assessment of the size distribution of broken rock at the extraction level drawpoints was carried out by Stratavision Pty Ltd. The block cave fragmentation (BCF) software program was used for this assessment. The program provides an indication of the primary fragmentation resulting from the initial caving processes and the secondary fragmentation that occurs when material higher up in the orebody is broken up as it is drawn down towards the drawpoints.

The work indicates that the New Afton deposit is highly fractured within all geotechnical domains and the proportion of the caved rock mass that will pass 2 m3 (a standard industry guideline) will be close to 100% for all areas, but with some potential for more coarse material during the early stages of caving in the hypogene material.

Sensitivity analysis of the fragmentation estimates to changes in fracture frequency in the potentially more competent hypogene domain in the lower portion of B3 indicates that a primary fragmentation of 77% <2 m³ is possible. The secondary fragmentation remains high with 98% <2 m³. Some large blocks may therefore be expected to report to the drawpoints in the early stages of production within this domain.

Only a small percentage of hang-ups and oversize rock (greater than 2 m3) reporting to the drawpoints are expected.

SURFACE SUBSIDENCE

Underground mining by caving methods will result in surface subsidence. Subsidence from massive underground three-dimensional orebodies is not well defined, the methods of analysis are not standardized, and the subsidence is normally discontinuous.

The impact of the subsidence above the New Afton deposit has been evaluated by both empirical and numerical methods. The empirical assessment of cave limits is only a crude analysis, but is sufficient as a preliminary guide to possible surface effects.

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The Crater Zone (cave limit; discontinuous deformations) developed using empirical methods is shown in Figure 19-5. The diagram also shows an exclusion zone (equivalent to the Cave Influence Zone) within which there is potential for (elastic, continuous) subsidence effects to occur. Permanent or difficult to replace surface infrastructure has been excluded from this area.

The numerical assessment used a novel design approach, combining the merits of draw simulation using Cave-Sim with a non-linear stress analysis of the ultimate cave geometry. However, the model is uncalibrated, and naturally such a model includes a number of important simplifying assumptions. The numerical analyses considered two variations on rock mass material properties, “Best” and “Worst”, but did not include discrete geological structures.

Figure 19-6 shows the extent of subsidence effects estimated using numerical modelling. The diagram shows two surface outlines resulting from the Abaqus modelling of the cave, termed “Best Credible” and “Worst Credible”, together with the surface outline resulting from the analysis of ‘underground mining on pit slope’ stability. Slope damage in the northern slopes occurs during the mining of B1. This zone grows during the subsequent mining of B3, and the failure zone extends along the Tertiary contact in the northern slope.

The subsidence crater and the influence zone will expand gradually as underground mining progresses. Monitoring of ground behaviour will follow the progression of underground and surface effects, and provide early indications of any unexpected behaviour, to allow contingency plans to be implemented. Automated pit slope monitoring is an ongoing exercise at the site.

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GROUND SUPPORT

Stress analyses were undertaken to determine the likely stability of the undercut and extraction levels for B1 and B2, where production for the initial years will take place, and where ground conditions are slightly poorer. Specific assessments for B3 were not carried out but will be required in the future, during a period of refinement of B3 layout and designs.

The support systems include 24 mm diameter full-column grouted rockbolts, cable bolts, F56 gauge weldmesh, mesh straps, and fibre-reinforced shotcrete. Two or three Henderson-style steel sets will be used in each drawpoint brow, and special measures are recommended to protect drawpoint pillar noses. The proportions of ground conditions in different categories were considered based upon statistical analyses of geotechnical data and the block model.

MAJOR EXCAVATIONS STABILITY

Stress analysis studies of the proposed crusher chamber demonstrated there is no interaction expected with the cave zones, and that the extent of yield around the crusher chamber will be limited to local re-entrants and brows. No other specific stability assessments were undertaken for other mine infrastructure.

MANAGEMENT PLAN AND MONITORING PROGRAM

A Ground Control Management Plan (GCMP) will be required. This is a ‘live’ document which captures all the key features of the operation’s ground control policies, practices, and procedures. The GCMP would include:

·	Details of the standard support designs.

·	An outline of the conditions under which variations from those standards are permitted.

·	The authorities / responsibilities for making any changes.

·	The support specifications in use at the mine.

·	A summary of the geotechnical monitoring programs in use at the mine.

A Geotechnical Monitoring Program will be an essential component of the Project’s rock mechanics program. The program will need to be established, operated, and maintained by a qualified rock mechanics engineer throughout the life of the Project. The objective

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of the monitoring program will be to provide data to enable the safe management of the mine in respect of the major mine hazards.

These hazards include but are not limited to:

·	Cave initiation, propagation, and surface impact/subsidence.

·	Inrush risk (mud, water, and airblast).

·	Major excavation stability (undercut and extraction levels, crusher).

·	Mine seismicity.

CAVE MONITORING

Monitoring of the cave is important and is planned to be done via borehole monitors and remote sensing. Draw control is also important and the draw must be managed. The recommended monitoring programs for both surface and underground are outlined below.

Surface monitoring should include the following:

·	Precise levelling across an array of monitoring stations distributed within and outside the predicted subsidence zone.

·	Open holes for monitoring of cave progression. These holes will be focused on the central and western areas of the deposit. A total of 12 holes are recommended.

·	Extensometers or time-domain reflectometry (TDR) cables also placed in surface boreholes drilled down to intersect the proposed cave at different elevations.

Underground monitoring should include the following:

·	‘Smart’ extensometers for monitoring ground movements on the undercut and extraction levels.

·	Additional absolute virgin stress measurements.

·	A Seismic Monitoring System.

A draw control plan should be designed and monitored, recording and managing the production from the drawpoints as well as recording and monitoring conditions at the drawpoints.

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MUDRUSH POTENTIAL

An assessment of the mudrush potential was completed for the 2007 Feasibility Study. There is a potential for mudrush hazard to develop during the life of the mine, due to the following:

·	Potentially mud-forming rock types exist within the cave material.

·	Potential for significant water inflow from terrestrial sources.

·	Potential for localized ingress of groundwater into the cave material along prominent structures and ungrouted drill holes.

·	Potential for ingress of wet mud and fines from the Afton pit as the cave breaks through the pit floor.

·	The generation of progressively increasing volumes of fines, due to comminution of the cave material as the cave matures.

The generation of fines within the Afton block cave is an unavoidable product of the caving process. As a result, it is recommended that engineering measures and procedures be developed, with the focus on minimizing the passage of water into the cave muck pile and on reducing the accumulation of fines within the cave muck pile by developing strict draw control strategies.

The following specific controls are recommended to minimize the potential for the development of mudrush hazards:

·	Develop and implement a draw management plan.

·	Conduct a drawpoint monitoring program to identify the nature of the material (particle size and moisture content).

·	Maximize use of remote-loaders.

·	Appoint a responsible person who is accountable for the ongoing task of assessing the mudrush potential.

·	Remove all free water from the New Afton open pit before mining commences.

·	Reduce the water content of the debris in the bottom of the pit such that the material is not subject to liquefaction when disturbed by subsidence or movement within the cave.

·	Minimize and control water inflow into the cave muck pile at every opportunity.

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The potential mudrush hazard has been based on limited information especially as related to the potential for fines generation within the cave muck pile. Additional rock property tests are recommended, including slake durability tests, for the specific rock types representative of those expected within the cave muck pile.

HYDROGEOLOGY

A hydrogeological evaluation was carried out to determine the present pit water balance and to predict likely inflows to the block cave. Total groundwater inflow to the pit/block cave is predicted to increase from about 4 L/s to 5 L/s under present pre-mining conditions, rising to about 8 L/s during block cave operations.

The highest transient inflows are predicted to occur once the back of the cave starts to fail and the zone of fracturing starts to migrate to the surface. Based on a six month period between the completion of the undercut and the initiation of surface subsidence, simulated flows ranged from 11 L/s to 17 L/s for the 0.5% and 1.0% porosity cases, with short term peak flows up to about 26 L/s. The transient flows are considered upper bound and will be re-assessed during initial mine developed and once the undercut development is initiated.

PIT DEWATERING AND DEBRIS STABILIZATION

Based upon a bathymetric survey and the final pit survey, it has been estimated that the pit contains approximately 2.6 Mm3 of supernatant water and 0.6 Mm3 of rock and soil debris. That volume was estimated to be increasing by about 0.15 Mm3 per year.

It is proposed to stabilize the debris material prior to block caving operations. The strategy involves removal of the supernatant water by pumping. Once the free water has been removed, a road will be constructed across the pit on top of the debris to access a series of up to 40 dewatering wells that will penetrate the full thickness of the debris to the base of the pit. Submersible pumps in each of the wells will continue to remove water until the moisture content of the debris has been reduced to approximately its liquid limit. Further consolidation will be achieved by either vacuum pumping from the existing wells or mechanical compaction.

The objective is to dewater to fines content of the debris, to approximately midway between its liquid limit and its plastic limit, so that a mudflow cannot occur. A period of

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approximately two years has been allowed for this process to be completed and verified, prior to commencement of caving in B1.

At the time of this report, the dewatering of the pit was underway with a forecast 280 days of pumping left to complete the initial dewatering. This would lead to completion of the first phase of dewatering near the end of the second quarter of 2010.

BLOCK CAVE DESIGN

UNDERCUT DESIGN

The undercut design is shown in Figure 19-7. The undercut will be mined from crosscuts driven at a nominal centreline spacing of 13 m. A low height undercut is proposed with a flat section over the drawbell and an “A” shape forming the crown of the major apex. A vertical interval of 17 m floor to floor has been adopted between the undercut level and the extraction level below. This distance is a reasonable compromise between the reduced stress effects on the extraction level achieved by increasing the interval and the practical issues associated with mining drawbells with a high level of consistency and reliability.

It is proposed that the undercut crosscuts will be mined 3.6 m wide by 4.0 m high. An access drive 5.5 m wide and 5.5 m high will be mined on the north side of the deposit. A slot drive will link the southern extremities of the crosscuts. The undercut crosscuts will be mined at a nominal gradient of 1:50 (2.0%) to the south. As the development will be driven down grade, face pumps will be necessary during the development stage. Drain holes down to the extraction level will be drilled where necessary. Each undercut crosscut will be ventilated using an auxiliary ventilation duct, both during development and when retreating the undercut.

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It is estimated that approximately 45% of the material blasted in the undercut will be extracted on the undercut level using 6-tonne to 7-tonne capacity load-haul-dump (LHD) units. This material will be tipped in ore passes leading to the ore transfer level. Five 2.4 m diameter passes from the B1 and B2 undercut level are proposed for this purpose. In order to maintain an overall undercut face angle of approximately 60° to the strike of the deposit, the cave face position in each crosscut will need to lag the face in the adjacent westerly crosscut by a minimum of 22.5 m.

The undercut will be drilled using 64 mm diameter holes at a ring spacing of 2.1 m. The inclination of the rings will be designed to reduce the lag between undercut face positions in adjacent crosscuts to a minimum. Approximately 93 m of drilling per ring will be required. A number of slot raises, up to 12 m long, will be required to initiate the undercut, a Cubex in-the-hole (ITH) drill rig with pneumatic hammers mounted on a Roger V-30 reaming head is proposed for this work.

DRAWBELL DESIGN AND CONSTRUCTION

Drawbell drives were designed at 4.0 m high x 3.6 m wide to provide adequate clearance for the drilling equipment. Drawbells will be constructed by drilling twin 760 mm pilot raises with a Cubex rig, and approximately 1,360 m of 64 mm blastholes with a hydraulic top hammer production drill. Pumped emulsion explosives and electronic detonators will be required to give the highest level of control and predictability over the blast. Drawpoint support work will be completed prior to firing the drawbell.

EXTRACTION LEVEL DEVELOPMENT

The extraction level crosscuts will be accessed from a 5.5 m wide x 5.5 m high access drive on the north side of the deposit. An offset herringbone layout is proposed, with the access crosscuts being 3.6 m wide by 3.6 m high, after application of surface support. These dimensions are suitable for the use of either diesel or electric LHD units.

Ore will be trammed back to short 2.4 m diameter ore passes connected to the production crosscut. Only one LHD will operate in a crosscut at any time and will only enter the access drive when moving from one production crosscut to the next, or when travelling to the workshop for maintenance. This arrangement avoids interaction with other LHDs or with development and ground support equipment that will need to use the access drive.

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To allow for the long life required, ground support standards will be high with extensive use of fibrecrete, rockbolts, and cables. Concrete roadways will be constructed in the extraction level crosscuts.

Intake air will enter the level from the access drive and flow past the drawpoints to exhaust ventilation raises at the southern end of the crosscuts leading to the ventilation level below. Ventilation raises will be fitted with small fans to control airflow.

Crosscuts will be graded down at 1:50 (2.0%), towards the ventilation raise such that water from the drawpoint dust suppression system, plus any groundwater emanating from the cave, flows to the raise and thus to the main collection sumps. It is proposed that water sprays installed at the drawpoints will be automatically controlled by vehicle movement sensors.

DRAWPOINTS

Drawpoints were designed to accommodate 6-tonne to 7-tonne capacity LHDs at 3.4 m wide by 3.4 m high. Drawpoints will be heavily supported with rockbolts, fibrecrete, and Henderson style sets. Figures 19-8 and 19-9 show a typical drawpoint arrangement. The tonnage drawn through each drawpoint will vary from 14,000 tonnes for those drawpoints whose purpose is only to extend the cave footprint to achieve the minimum span required for caving, up to 165,000 tonnes. Concrete roadways will be constructed in all drawpoints, with steel wear rails embedded in the concrete in those drawpoints where in excess of 75,000 tonnes is planned to be extracted. Drawpoints will be fitted with automatic dust suppression sprays, activated when a loader enters the drawpoint.

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ORE AND WASTE HANDLING

As part of the Feasibility Study, a trade-off study to assess the merits of using either shaft or conveyor haulage for ore transfer to surface was completed. Conveyor handling was chosen as a lower overall cost alternative.

ORE AND WASTE HANDLING REQUIREMENTS

During the initial development of the mine, all ore and waste will be trucked to the surface. Following commissioning of the underground crushing and conveying system, truck haulage in the conveyor decline will be limited to delivery of bulk materials to the mining areas and possibly some waste transfer to surface.

Approximately 12% of the ore to be mined will originate from B3, located approximately 120 m below the crusher feed position. Ore from this area will be mined later in the mine life and will be trucked up to the crusher via a decline.

A summary of the estimated total quantity of ore and waste mined over the life of the Project is shown in Table 19-1 and Figure 19-10.

Prior to installation of the underground crusher and conveyors, ore and waste will be trucked to the surface using a fleet of 50 tonne capacity articulated haul trucks. Trucking routes out of the mine will include the conveyor ramp, except during installation of the conveyor, and the existing exploration ramp. Either route could be used for ore and waste but, in general, ore will be trucked via the conveyor ramp to a stockpile adjacent to the mill and waste, via the exploration ramp for disposal in the open pit. It is envisaged that waste hauled to the surface via the conveyor ramp will be placed on top of the future subsidence area.

The estimated total average annual ore and waste movement during the initial construction period is shown in Figure 19-10. Material movement requirements will peak at approximately 2,400 tonnes per day prior to crusher commissioning (Q2 2012). The material movement increases to 4,900 tonnes per day when the crusher is commissioned and ramps up to a peak of 11,900 tonnes per day.

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TABLE 19-1 LOM ORE AND WASTE TONNAGE SUMMARY

New Gold Inc. – New Afton Project

	B1 Ore	B2 Ore	B3 Ore	Produc-tion	Ore Develop-ment	Ore Mined	Waste	Total Mined
	000 t	000 t	000 t	000 t	000 t	000 t	000 t	000 t
2009	-	-	-	-	-		93	93
2010	-	-	-	-	-	-	323	323
2011	-	108	-	108	97	205	393	598
2012	4	1,933	-	1,937	85	2,022	203	2,225
2013	812	3,188	-	4,000	85	4,085	131	4,216
2014	2,155	1,845	-	4,000	29	4,029	46	4,075
2015	2,625	1,375	-	4,000	3	4,003	17	4,020
2016	2,907	1,093	-	4,000	-	4,000	-	4,000
2017	2,455	1,545	-	4,000	-	4,000	-	4,000
2018	2,232	1,768	-	4,000	-	4,000	7	4,007
2019	2,744	1,256	-	4,000	34	4,034	135	4,170
2020	3,283	424	293	4,000	28	4,028	52	4,080
2021	2,009	250	1,740	4,000	-	4,000	-	4,000
2022	1,956	244	1,800	4,000	-	4,000	-	4,000
2023	1,956	244	1,800	4,000	-	4,000	-	4,000
2024	687	86	227	1,000	-	1,000	-	1,000
				-
Total	25,826	15,358	5,861	47,045	361	47,406	1,402	48,808

FIGURE 19-10 ORE AND WASTE TONNAGE

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CONVEYOR DECLINE

The conveyor decline has been designed at a gradient of 1:6 (16.7%) with four legs, varying in length from 700 m to 1,200 m. Transfer stations were designed at the head and tail end of each leg. The cross-sectional dimensions, at 5.5 m wide by 6.5 m high, provide adequate clearance for a loaded 50-tonne capacity truck and the conveyor structure. A typical cross-section through the conveyor decline is shown in Figure 19-12.

The decline is being developed using drill and blast techniques with diesel powered trucks and large capacity LHDs. Ground support for the decline includes the use of rock bolts and fibrecrete.

It is proposed that the conveyor structures would be of modular construction and suspended from chains attached to the back of the decline. The following belt specification was used in the Feasibility Study:

·	1,050 mm wide ST3150

·	13 mm top cover, abrasion, cut, gouge and fire resistant 7 mm bottom cover

·	fabric breaker cover above cords

·	rip detection loops at 50 m spacing

·	fire detection systems installed along the length of the belt with automatic water sprays installed in head and tail areas

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CONVEYOR TRANSFER ARRANGEMENTS

The design philosophy for the transfer stations is to minimize the excavation volume required; to keep maximum excavation spans below 10 m; to simplify installation, maintenance and belt replacement activity; and to provide belt replacement facilities at the head end using a 4.5 m high roll of belting.

The geometry of the conveyor decline results in shallow angles at the transfer points. A transfer arrangement having separate head end and tail end chambers with a short interconnecting ore bin was selected on the basis of its relative simplicity and lower cost.

With the exception of the portal conveyor, each conveyor will be driven from the head end using dual drives. The head end drive floor will be elevated, providing a roadway beneath the floor for maintenance vehicle access to the head end chamber and belt reeling station.

A tail end belt take-up arrangement was selected for all transfer stations as it provides the best combination of minimizing excavation spans and overall excavation volume. The conveyor decline and transfer stations were designed such that trucks and other vehicles using the decline do not have to pass through the transfer chambers.

INSTALLATION OF CONVEYOR SYSTEM

At various stages during the installation of the conveyor system, it will be necessary to suspend trucking activities in the conveyor decline. During these periods, ore will be trucked to surface via the alternative route through the open pit portal and the open pit access road to the mill stockpile.

To speed the installation process, it is envisaged that idler frames in 6 m lengths complete with idlers will be pre-assembled on surface and delivered down the decline and installed using a suitably modified scissor-lift truck. Laser survey control will be used to align the holes for the conveyor suspension bolts and the idler frames.

ORE TRANSFER FROM THE EXTRACTION LEVEL TO THE CRUSHER

LHDs operating on the extraction level will feed ore to short ore passes leading to the ore transfer level located approximately 17 m below. No grizzlies are proposed at the

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top of the passes, which will be sized to accept the largest possible rock capable of being delivered by the 6-tonne to 7-tonne capacity LHDs.

Two 17-tonne capacity LHDs will load ore from the base of the ore passes and tram it to the crusher dump pocket. In order to protect the ore pass brows from excessive wear, steel sets embedded in high strength fibrecrete similar to those proposed for the production drawpoints will be installed at the base of each pass. Brow chains will also be fitted at the bottom of the passes. The passes will be fitted with level sensors and lights on both the extraction and transfer levels to ensure the brow of the pass is always kept closed.

Dual tramming routes were designed on the transfer level to avoid interaction between haulage units. There is the potential for fully automated operation of the loaders on this level.

The transfer drives were designed at 5.5 m wide by 4.0 m high and will be mined at a nominal gradient of 1:40 (2.5%). A high strength concrete road bed will be installed along with embedded steel wear bars in the crosscuts leading to the base of each pass.

CRUSHER

A 150-tonne to 600-tonne capacity crusher dump pocket designed to allow two-sided tipping by LHDs and single side tipping by 50-tonne capacity trucks will be installed above a 2.4 m wide apron feeder. The feeder will control the feed rate to a 2,000 mm x 1,500 mm nominal size single toggle jaw crusher. A fixed rock breaker installed at the crusher feed hopper will minimize production delays due to crusher bridges.

The crushed material is discharged into a pocket and bin of approximately 200-tonne capacity located directly below the crusher outlet cavity. The crushed ore bin discharges onto the picking conveyor belt, where a tramp metal detection and collection system will remove metal objects before ore is fed to the main decline belt system.

Ore from B3 will be trucked from a truck loading bay near the extraction drift directly to the crusher dump pocket or to a short pass adjacent to the crusher. A maximum production rate of approximately 1.8 Mtpa has been scheduled from B3.

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WASTE HANDLING

It is envisaged that waste mined after commissioning of the crusher and conveyor system will be trucked to surface via the conveyor decline, however, provision has also been made for waste to be batched through the system on an as-needs basis. The short ore pass adjacent to the crusher designed to handle ore from B3 will be used as a waste storage pass. Very little waste will be mined once ore production commences from B3.

The transfer tower at the head end of the portal conveyor will be fitted with a head chute and a two-way transfer to either the mill feed conveyor or waste conveyor.

PRE-PRODUCTION SCHEDULE

Exploration of the New Afton deposit included an underground drive and access from the 512 m elevation of the open pit. This development provided a starting point for the mine development coupled with a new portal for the conveyor gallery collared near the mill. Mine development commenced in 2007 and continued until November 2008, when it was halted to conserve cash in the difficult financial markets. The mine development in 2007 and 2008 was carried out by a mining contractor, Cementation Inc.

Mine development restarted in February 2009 as New Afton changed from contractor mining to owner mining. The mine development is advancing the conveyor galleries and the main access. Initial ore from development, undercutting, and production drawpoints will be stockpiled in a temporary stockpile area adjacent to the mill. Milling operations were scheduled to commence in month 37 and will increase to the planned production rate of 4 Mtpa over a three-month period. Operations are proposed to continue at the planned rate for 11 years.

Lateral development scheduling has been completed using the following assumptions:

·	All remuck bays have been designed and included in the estimate of total drift metres.

·	An allowance for overbreak of 10% of the designed excavation volume is added when estimating tonnages of waste from development.

·	Waste is assigned a density of 2.6 t/m³. The density of development ore is taken from the resource block model.

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·	The advance of any single heading is limited to a maximum of 30 m/week. In the case of the extraction level crosscuts and undercut development, this is reduced to 20 m/week.

·	The maximum drawpoint development rate is scheduled at 10 m/week.

·	In multiheading development situations, the maximum development rate scheduled per development jumbo is limited to 50 m/week.

·	Development of the extraction level crosscuts and drawpoints are scheduled in two stages. Initially, the crosscut is advanced to its southern extremity. A minimum period of four weeks is scheduled to pour and cure the concrete roadway before development of the first drawpoint.

·	Construction of drawpoints and drawbell crosscuts is scheduled after the roadway is complete and the undercut has passed overhead.

·	All bored raises are scheduled at an average advance rate (including rig up, piloting, reaming, and rig down activities) of 30 m/week.

·	All distances are based on the centreline to centreline distance. The exception to this is the extraction level drawpoints where the length is measured from the wall of the extraction crosscuts.

PROGRESS TO SEPTEMBER 30, 2009

Actual development by Company crews for the period February to September 2009 totalled 1,072 m vs. the budget of 1,490 m. The shortfall in advance reflects the development and training of new mining crews and limited heading availability. Management indicates that this development shortfall is being addressed by moving from 5 day per week 10.5 hour shifts to 7 day per week 11 hour shifts. Development rates will also benefit from the availability of close-spaced multiple headings in 2010.

Productivity rates used in the LOM reflect actual productivity to date. Development cost to date is $6,829/m to the end of September based on mining and maintenance costs, but excluding indirects, such as general and administration (G&A) costs. Comparable LOM development costs are $5,733/m. The LOM cost estimate reflects increasing expertise within the crews, increased efficiencies resulting from 24:7 operation and longer shifts, and multiple close-spaced headings. Actual productivity to date was taken into account in estimating advance rates for the LOM.

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INFRASTRUCTURE

MINE ACCESS

The deposit is bounded on the southern side by a major fault structure (the Hanging Wall Fault and Picrite), which is considered to be difficult to cross; therefore, the permanent access and underground infrastructure has been designed on the north (footwall) side of the deposit.

The existing exploration portal (located within the open pit) and decline provided early access for the undercut and extraction level development and a conveyor decline portal has been developed close to the proposed mill site at an elevation of 674 m.

After the commencement of production, caving will cause access through part of the exploration decline to be lost. Prior to this occurring, a service raise (VR5), equipped with an emergency hoisting system will be installed to provide a separate connection between the mine workings and surface. During the remaining life of the mine, the main access will be via the conveyor decline, with the service raise and hoisting system providing a second means of egress for use in an emergency.

MINE VENTILATION

The primary ventilation circuit will be a push-pull system with six main surface fans installed on three intake shafts and three exhaust shafts. The conveyor decline will be an exhaust airway. The total required airflow has been estimated at 500 m3/s. Exhaust raises will be 3.5 m diameter and will each have an airflow of approximately 140 m3/s, with an additional 80 m3/s returning via the conveyor decline. The three intake raises will be 3.5 m diameter and will each have an airflow of approximately 167 m3/s.

Flow rates and pressures in the proposed ventilation circuit were modelled using Ventsim software. Six horizontally mounted axial flow fans are proposed to meet the required duty. A number of secondary and booster fan arrangements within the mine will ensure distribution of air to all working areas.

Simple, robust and automated circuit control devices are proposed to create an efficient ventilation system. Automatically activated water sprays will be fitted to production drawpoints on the extraction level to minimize dust. Dust extraction systems will be

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installed on orepasses, at conveyor transfer points and in the crusher station.

AMC has estimated an average surface rock temperature at New Afton of approximately 12°C and a geothermal gradient of approximately 3°C per 100 m depth. There is some potential for hot working conditions to be encountered in the deeper parts of the mine while intensive development is being undertaken (particularly during summer), but these situations can be avoided provided that secondary ventilation systems are properly designed, installed, and maintained.

Between 1978 and 1993, approximately 23% of days at New Afton had average temperatures below freezing. To avoid the problems associated with air intake temperatures consistently below freezing, it is proposed to install natural gas, direct fired heating plants on each of the intakes. These plants will be installed near the portals with warm air ducted to the portal entrances and the intake raises. Only a portion of the air entering the portals will be heated. The quantity and temperature of the heated air will be adjusted to maintain the overall temperature of air entering the mine above freezing.

UNDERGROUND SHOPS AND SERVICES

An underground workshop will be constructed adjacent to the extraction level for the mobile equipment fleet maintenance. An underground fuel bay will be installed with a dedicated return airway and a fire suppression system. An underground magazine has been designed to hold approximately one week’s consumption of explosives, with supply from surface through regular visits from explosives suppliers.

Lunch room and refuge station facilities will also be installed on the extraction level.

MINE DEWATERING

The underground pumping system is designed as a staged system with an installed pumping capacity of 50 L/s. The main pumps will be remotely monitored and controlled.

COMPRESSED AIR AND ELECTRICITY

A mine-wide compressed air system is not planned for the entire mine. A local system will be installed in the workshop and crusher area for maintenance and other miscellaneous activities. Compressed air for other activities will be provided by on-board and mobile compressors. Electrical power will be reticulated through the mine at

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13.8 kV via a ring main system.

COMMUNICATIONS

The proposed mine communications system includes telephone communications (voice over internet), a leaky feeder radio network, an asset tracking system, video monitoring systems and programmable logic control systems for pumps, fans and ventilation control systems, ore pass level monitoring, crushing and conveying systems.

MINE EQUIPMENT

The current mine equipment fleet is in service for the development of the mine access. As the undercut development proceeds, there will be additional equipment required for the long hole drilling and then for mine production. The current mine equipment fleet is shown in Table 19-2.

TABLE 19-2 UNDERGROUND MOBILE EQUIPMENT
New Gold Inc. – New Afton Project

Type	Manufacturer	Model	Number	Capacity
LHD	Atlas Copco	ST-6C	1
LHD	Atlas Copco	ST-1520	3
Truck	Atlas Copco	MT-5010	3
Truck	Atlas Copco	MT-444	1
Jumbo	Atlas Copco	Boomer M2C	2
		Boomer 322	1
Bolter	Atlas Copco	Bolter 358	1
Bolter	Atlas Copco	Bolter MC	1
Boom truck		M40	1
Scissor Lift		HL 6000	3
Utility	Toyota	various	10
Sprayer		7110	1
Sprayer		6050W	1
Sprayer		6050EW	1
Transmixer		M40	2
Transmixer		Utimec 1600	1
Transmixer		Utimec 1500	1
Dozer	Komatsu	D65EX-15	1
Loader	Caterpillar	930G	2
Grader	Caterpillar	140M	1
Tractor	Kubota	T100	1

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The planned capital period additions to the mine equipment fleet are shown in Table 19-3.

TABLE 19-3 CAPITAL EQUIPMENT FOR PRODUCTION

New Gold Inc. – New Afton Project

Type		Number
Development Jumbos (small)		3
Development Bolters (small)		3
Cable Bolter		1
Transmixers		3
Shotcrete Sprayers		2
Trucks		2
Toyotas		4
Boom Truck		1
Integrated Toolhandler		2
Long Hole Drill		3
Cubex Drill		1
Development LHDs Small		6
Production LHDs Large		2
Production LHDs Small		6
Cassette Truck		1

Over the life of the mine, there will be equipment replacement as indicated in Table 19-4. The costs for these items are included in sustaining capital.

TABLE 19-4 SUSTAINING CAPITAL U/G EQUIPMENT

New Gold Inc. – New Afton Project

Item	Qty
Production LHDs - Large	3
Development LHDs Diesel - Small	3
Production LHDs Electrical - Small	9
Trucks	1
Forklift Integrated Toolhandler	6
Long Hole Drill – Solo	1
Shotcrete Sprayers	1
Shotcrete Haulers (Two per sprayer)	3
Scissor Truck	1
Flatbed Truck	2
Crane Truck	1
Light Vehicles	35
Support Equipment	3

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MINERAL PROCESSING

Proposed process and recoverability are described under Section 16.

LIFE OF MINE PLAN

The Life of Mine Plan (LOMP) in this report commences on July 1, 2009, with the continuation of the mine development followed by the mill construction and then the production phase of the LOMP. Activity and expenditures prior to July 1, 2009, are considered sunk.

Key mine development and production estimates for the preproduction period are outlined in Table 19-5. The table covers the period ending June 2012 after which the mine is forecast to be in production.

TABLE 19-5 LOMP DEVELOPMENT

New Gold Inc. – New Afton Project

Month	Development Metres	Bulk (m3)	Vertical	Waste (t)	Ore (t)
Jul-09	131	-	-	10,394	-
Aug-09	166	-	-	13,164	-
Sep-09	173	648	-	15,250	-
Oct-09	174	720	-	15,535	-
Nov-09	229	540	-	19,438	-
Dec-09	229	540	-	19,438	-
Jan-10	258	1,548	-	24,252	-
Feb-10	247	1,260	28	23,557	-
Mar-10	262	1,836	180	30,234	-
Apr-10	262	1,404	92	26,775	-
May-10	262	1,836	130	28,913	-
Jun-10	262	1,836	-	26,230	-
Jul-10	262	1,836	-	26,361	-
Aug-10	290	900	-	26,126	-
Sep-10	317	-	-	25,957	-
Oct-10	317	-	80	28,108	-
Nov-10	316	-	111	28,803	-
Dec-10	315	-	110	27,848	-
Jan-11	293	-	149	18,096	4,241
Feb-11	477	147	120	30,581	4,241
Mar-11	724	196	120	35,549	8,856
Apr-11	675	2,900	150	39,815	8,856

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Month	Development Metres	Bulk (m3)	Vertical	Waste (t)	Ore (t)
May-11	674	2,900	120	35,803	8,856
Jun-11	675	2,900	120	35,803	8,856
Jul-11	695	1,944	120	34,799	17,018
Aug-11	684	1,944	132	33,137	17,018
Sep-11	696	1,704	120	33,137	17,018
Oct-11	735	-	153	31,559	36,703
Nov-11	741	-	166	32,296	36,703
Dec-11	727	-	140	32,333	36,703
Jan-12	643	-	140	30,521	42,075
Feb-12	566	-	153	23,808	42,075
Mar-12	522	-	37	17,079	42,075
Apr-12	471	-	20	13,081	132,249
May-12	471	-	26	13,424	132,249
Jun-12	471	-	45	13,613	132,249

The annual production and development activity is summarized in Table 19-6. This table commences in July 2009 and is for annual periods, with only six months in 2009.

TABLE 19-6 LOMP DEVELOPMENT TONNAGE

New Gold Inc. – New Afton Project

Year	Development Metres	Bulk (m3)	Vertical	Waste (t)	Ore (t)
2009	1,101	2,448	-	93,220	-
2010	3,371	12,456	731	323,166	205,071
2011	7,799	14,635	1,610	392,907	2,021,904
2012	6,091	-	542	203,216	4,084,882
2013	5,126	-	179	131,482	4,029,197
2014	2,034	-	80	45,628	4,002,627
2016	580		-	17,420	4,000,029
2017	-		-	-	3,999,979
2018	-		-	-	3,999,965
2019	93		-	6,936	4,034,421
2020	2,872		219	135,450	4,027,975
2021	2,221		37	52,430	4,000,000
2022	-				4,000,000
2023	-				4,000,000
	-				999,690
Totals	31,288	29,539	3,398	1,401,856	47,405,739

The LOMP mill production estimates are shown in Table 19-7.

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TABLE 19-7 LOM MILL FEED
New Gold Inc. – New Afton Project

Year	Hypogene (kt)	Mesogene (kt)	Supergene (kt)	Ore Milled (kt)	Cu (%)	Au (g/t)	Ag (g/t)	Hg (ppb)	As (ppm)
2012	2,217	10	-	2,227	0.84	0.74	1.47	412	131
2013	4,011	74	-	4,085	0.86	0.72	1.39	503	162
2014	3,767	262	-	4,029	0.82	0.69	1.66	582	159
2015	3,465	538	-	4,003	0.91	0.77	2.14	886	234
2016	3,242	758	-	4,000	1.07	0.81	2.66	1,330	415
2017	3,327	673	0	4,000	1.17	0.83	2.59	1,150	406
2018	3,423	577	0	4,000	1.12	0.75	2.42	1,057	393
2019	3,363	655	17	4,034	1.14	0.66	2.74	1,248	499
2020	3,207	533	288	4,028	1.04	0.59	2.45	1,236	427
2021	3,406	178	416	4,000	0.95	0.67	1.90	760	205
2022	3,390	98	511	4,000	0.87	0.64	1.71	818	218
2023	3,395	53	553	4,000	0.63	0.46	1.25	751	188
	791	12	197	1,000	0.50	0.36	1.04	757	142
Total	41,003	4,421	1,982	47,406	0.95	0.69	2.03	909	289

The recovery of metals and the concentrate production is shown in Tables 19-8 and 19-9.

TABLE 19-8 LOM METAL RECOVERY
New Gold Inc. – New Afton Project

Year	Cu (%)	Au (%)	Ag (%)
2012	76.34	75.47	66.52
2013	89.68	88.37	77.06
2014	88.99	87.55	77.63
2015	89.28	87.25	75.81
2016	90.00	87.85	75.12
2017	90.70	88.34	73.92
2018	90.57	88.55	75.26
2019	90.43	88.70	76.65
2020	87.83	87.09	74.85
2021	86.36	86.93	73.84
2022	84.83	85.94	74.92
2023	81.07	83.56	78.29
2024	74.96	79.12	75.44
Total	87.82	86.74	75.27

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The average recovery of mercury and arsenic to the concentrate is estimated to be 70.7% and 67.6%, respectively. Recovery for copper, gold, and silver was estimated as outlined in Section 16, plus a factor of 87% was applied to 2012 recoveries to account for start-up uncertainty.

TABLE 19-9 CONCENTRATE PRODUCTION

New Gold Inc. – New Afton Project

Year	Con (dmt)	Con (wmt)	Cu (%)	Au (g/t)	Ag (g/t)	Hg (ppm)	As (%)
2012	61,500	66,848	23.24	20.11	35.42	9,281	2,441
2013	113,071	122,903	28.00	23.00	38.66	12,946	3,507
2014	104,775	113,885	28.00	23.36	49.44	15,767	3,702
2015	116,610	126,750	28.00	23.17	55.70	20,462	5,009
2016	137,390	149,337	28.00	20.73	58.27	26,245	7,964
2017	151,468	164,639	28.00	19.48	50.49	20,651	7,395
2018	145,346	157,985	28.00	18.25	50.05	20,409	7,695
2019	147,879	160,738	28.08	15.94	57.37	24,671	10,042
2020	126,765	137,788	29.07	16.37	58.34	26,516	8,707
2021	111,502	121,198	29.30	21.03	50.24	17,552	4,223
2022	98,664	107,243	29.93	22.24	52.02	21,686	5,042
2023	65,542	71,241	31.06	23.55	59.58	31,706	6,782
2024	11,381	12,370	32.94	25.10	68.82	44,123	7,122
Total	1,391,892	1,512,927	28.32	20.29	52.05	21,165	6,419

MARKETS

The New Afton mine’s principal commodities are copper concentrates containing copper and gold and doré containing gold. The principal commodities are freely traded at prices that are widely known so that the prospects for the sale of production are virtually assured.

Copper is a principal metal traded on the London Metal Exchange (LME) and has total price transparency. Prices are quoted on the LME for Copper Grade A and can be found at www.lme.com. The average copper price for the first nine months of 2009 was $2.11 per pound. Current prices as of September 30, 2009, are $2.81 per pound. The three-year and five-year rolling average prices through the end of September 2009 are $2.92 and $2.61 per pound, respectively. This Technical Report uses the long-term

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price forecasts from the major banks with a long-term average price for copper of $2.00 per pound for the economic analysis.

Gold is a principal metal traded at spot prices for immediate delivery. The market for gold trading typically spans 24 hours a day within multiple locations around the world (such as New York, London, Zurich, Sydney, Tokyo, Hong Kong, and Dubai). Daily prices are quoted on the New York spot market and can be found on www.kitco.com. The average New York spot gold price for the first nine months of 2009 was $930 per troy ounce. The current New York price, as of September 30, 2009, is $997 per troy ounce. The three-year and five-year rolling average prices through the end of September 2009 are $807 and $684 per troy ounce, respectively. This Technical Report uses the long-term price forecasts from the major banks with a long-term average price for gold of $800 per troy ounce for the economic analysis.

Silver trading follows a pattern that is similar to that of gold (as described above). Daily prices are quoted on the New York spot market and can be found on www.kitcosilver.com. The average New York silver price for the first nine months of 2009 was $13.68 per troy ounce. The current New York price, as of September 30, 2009, is $16.39 per troy ounce. The three-year and five-year rolling average prices through the end of September 2009 are $13.94 and $11.87 per troy ounce, respectively. This Technical Report uses the long-term price forecasts from the major banks with a long-term average price for silver of $12.00 per troy ounce for the economic analysis.

For the base case Scott Wilson RPA used metal prices as shown in Table 19-10.

TABLE 19-10 METAL PRICE AND EXCHANGE RATE ASSUMPTIONS

New Gold Inc. – New Afton Project

Metal	2009	2010	2011	2012	2013	Long Term	Reserves	Resources
Gold (US$/oz)	940.00	965.00	950.00	900.00	850.00	800.00	800.00	900.00
Silver (US$/oz)	13.75	15.00	14.00	13.25	13.00	13.00	12.00	15.00
Copper (US$/lb)	2.00	2.40	2.50	2.50	2.40	2.00	2.00	2.50

C$/US$	0.90	0.90	0.90	0.90	0.90	0.85	0.85	0.85
US$/C$	1.11	1.11	1.11	1.11	1.11	1.18	1.18	1.18

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CONTRACTS

New Afton does not have contracts in place for mining, construction, Engineering, Procurement and Construction Management (EPCM), operating supplies and reagents, concentrate shipping or concentrate treatment. Scott Wilson RPA expects that all of these items can be organized in a timely manner before operations commence. New Afton does not plan to engage mining contractors for the underground development or production.

ENVIRONMENTAL CONSIDERATIONS

The New Afton Project lies within the Interior Plateau subdivision of south-central British Columbia. The area’s complex geology, glacial history and continental, semi-arid climate have given rise to many regional landscapes. The Project area lies at about 720 MASL and drains into Kamloops Lake, the second-largest water body contained in the Thompson River watershed, a major tributary of the Fraser River. Much of the Project area has been greatly altered such that natural undulating grasslands formed on till-covered drumlinoid terrain and interspersed by numerous small, alkaline water bodies have been replaced by level rock dumps, open pits, and severely compacted borrow areas. The dominant water bodies are now related to mining-related impoundments. The baseline environmental studies undertaken considered both the natural and anthropogenic landscapes. The two primary landscape types are:

·	Extensively disturbed areas, reconfigured and/or reclaimed as a result of historic mining activities and infrastructure development; and

·	Undisturbed landforms, which are similar to those that existed prior to the development of the New Afton mine, consist of Ponderosa Pine–bunchgrass vegetation interspersed by discontinuous alkaline water bodies and ephemeral drainage altered by grazing land use.

Environmental studies undertaken for the Project included:

a)	Climate studies, air quality, meteorology

Noise

c)	Emissions analysis

d)	Surficial geology, topography, terrain and soils

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e)	Surface water quality and quantity

f)	Groundwater quality and quantity

g)	Hydrological modelling for water balance management

h)	Pit Lake bathymetry

i)	Aquatic resources including fish and fish habitat, wetlands, primary and secondary producers, physical limnology

j)	Terrestrial ecosystems and vegetation

k)	Invasive plants

l)	Wildlife, including species at risk

m)	Land status and use

n)	Land capability

o)	Archaeology

p)	Metal leaching and Acid Rock Drainage status and potential

Environmental baseline studies were undertaken on the Project site between April and December 2006. All baseline studies were implemented in accordance with provincial data collection guidelines where applicable, and generally accepted scientific procedures. The objectives and results of baseline studies were reviewed with the Kamloops Indian Band and Skeetchestn Indian Band in support of project-related consultations.

PERMITTING

In addition to the Mine Permit, a number of permits, consents, and authorizations are required to develop the Project. Table 19-11 sets out the permits which have been obtained, permits required by the Province of British Columbia, and permits required by the Government of Canada.

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TABLE 19-11 AUTHORIZATIONS AND LICENCES

New Gold Inc. – New Afton Project

Government Authorizations, Licences and Permits obtained by New Gold

Act (Authority)	Permit	Purpose
Mines Act (MEMPR)	Mines Act Permit MX 3-217	Issued to the Company
		Authorizes mineral exploration subject to provisions of Mine Health, Safety and Reclamation Code for Mines S(10).
Mineral Tenure Act (MEMPR)	Mineral Claims	Issued to the Company
		Conveys title to subsurface mineral rights. Annual fee as work commitment or cash in lieu to maintain in good standing. Issued to the Company.
	Mining Lease – Tenure Number 546063
Transportation Act (MOTH)	Permit / Licence	1. Authorization to construct a High Voltage (“HV”) line crossing of the Trans-Canada Highway. Will be applied for by B.C. Hydro prior to construction of HV line. This may no longer be required with a potential modification to the alignment of the connecting HV line to the minesite substation.
		2. Legal access agreements required over certain lands – Pending;
		3. Industrial access permit for the access easement to Afton Mine Road. – Pending.
		4. Access permit for the existing service road (water pipeline access) to Trans-Canada Highway.
Environmental Management Act (MOE)	Waste Management Permit (Effluent)	Application by Company pending. Authorizes introduction of waste into the environment (tailings, waste rock), subject to conditions, regulations or Codes of practice.
	Waste Management Permit (Sewage)	Application by Company pending (will be included in Effluent Application).
		Authorizes discharge of sewage into the environment, subject to conditions.
		Authorizes solid waste storage and disposal subject to conditions.
	Waste Management Permit (Air Emissions)	Pre-Application meeting February 14, 2008. Application by Company pending.
		Authorizes emissions to the air, subject to conditions.
Water Act (MOE)	Licence	Application by Company pending.
	(Notice of Intention)	Authorizes licensing, storage, diversion and use of water.
	(Diversion)	Required to secure water withdrawal, e.g. Kamloops Lake, develop diversion or impoundment structures on lease.
	(Use)	Issued by Comptroller of Water Rights.
Wildlife Act (MOE)	Permit (to relocate any wildlife)	Applicability subject to seasonal requirements.
		Authorizes removal of wildlife to another approved site by qualified individual. Requirement to be assessed for listed amphibians in Pothook Pit, Waste Dump Pond and NW Emergency Pond prior to works.

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Act (Authority)	Permit	Purpose
		No permit is required at present. Requirement to be reviewed prior to construction commencing.
Forest Act (Ministry of Forests)	Licence to Cut	Free Use Permit issued October, 2007.
Forest Act (Ministry of Forests)	Licence to Cut	Authorizes cutting of Crown timber not exceeding 50 cubic metres
Pipeline Act (OGC)	Permit – Pipeline Regulation	Application by Pembina Pipelines to OGC on behalf of Company – not immediately required.
Pipeline Act (OGC)	Permit – Pipeline Regulation	Authorization of work to relocate a section of oil pipeline north of the Afton Pit outside of subsidence zone. Status – pending.
Health Act (MOH – Interior Health Inspector) or Environmental Management Act (MOE)	Industrial Camps Health Regulation Permit or Camp Construction and Operating Permit – Applicable permit will be for water treatment plant.	Application by Company pending. Establishes health and sanitary conditions as per scale of camp and nature of facilities on site. Tied in with SRMDRC 60 - day review and inspection during construction phase.
		Approval for source and waterworks(domestic water treatment facility) and issuance of operating permit
Fire Services Act	Authorization	Fuel storage approval; routine, may be subject to inspection.
Agricultural Land Commission Act	Application	Agricultural Land Commission approved nonfarm use of Agricultural Land Reserve lands in October, 2007.
Terasen Gas	Permit	Application by B.C. Hydro on behalf of Company pending.
Terasen Gas	Permit	Permit to authorize crossing of natural gas pipeline with overhead HV electrical power line (138 kV).
Integrated Pest Management Act (MOE)	Permit pursuant to S(6)	Application by Company pending. Provides authorization to apply chemical treatments for control of noxious weeds.
Explosives Act R.S. C-E15	Licence	To authorize an explosives storage magazine.
	Licence (Explosives Factory)	To authorize the manufacture of explosives on site.
	Permit	To authorize vehicles used for transportation of explosives.
Radio Communication Act	Radio Licences	To establish and operate radio frequencies and related infrastructure.
Atomic Energy Control Act	Radio-Isotope Licence	To authorize nuclear devices, such as slurry density flowmeters.
Fisheries Act	Section 35 – authorization of works in and about fish habitat.	Permit requirement assessed for water pipeline intake maintenance. Department of Fisheries and Oceans approved conceptual design of works and will issue applicable authorizations when works are to commence.

Notes: MEMPR – Ministry of Energy, Mines and Petroleum Resources; MOTH – Ministry of Transportation and Highways; MOE – Ministry of Environment; OGC – Oil and gas Commission; MOH – Ministry of Health

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SOCIOECONOMIC AGREEMENTS

Two first nations, the Kamloops Indian Band and the Skeetchestn Indian Band (collectively, the “Bands”) have asserted aboriginal rights and interests in an area comprising the New Afton Project. Discussions with these Bands commenced in February 2006. After a number of meetings, information exchanges and discussions, the Bands and New Gold signed letters of intent in September 2006 which provided among other things for an exchange of information and continued dialogue between the parties. Further discussions with the Bands during 2007 and 2008 culminated in the execution and delivery of a formal agreement between the Bands and New Gold dated March 20, 2008 (the “Participation Agreement”).

One of the principal objectives of the Participation Agreement is to establish a cooperative and mutually beneficial working relationship between the Bands and New Gold with respect to the development and operation of the New Afton Project. The Participation Agreement provides that the Bands consent to the New Afton Project and agree not to challenge before a court of law any New Gold interests or permits related to the Project. New Gold undertakes to provide the Bands with certain economic and social benefits including preferential hiring, education, training and business opportunities.

The Participation Agreement contemplates that New Gold will pay to a socioeconomic trust to be created for the benefit of Band members an amount of $250,000 in 2009. Thereafter, each year in which commercial production occurs at the mine, an amount of 0.5% NSR, or a minimum of $250,000 a year, shall be paid to the trust, with the minimum payment to be increased to $500,000 for a given year if copper prices in that year exceed $2.50 a pound.

After the costs of developing, permitting, constructing, and equipping the mine have been repaid to New Gold, the payment to the trust will increase to the greater of $1,000,000 a year, or 1% of NSR in that year. Should copper prices exceed $2.50 a pound in a given year, this payment will increase to 1.5% of NSR for that year. Should copper prices exceed $3.00 a pound in a given year, this payment will increase to 2% of NSR for that year.

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The Participation Agreement also grants the Bands a right of first refusal to purchase that portion of the Afton Surface Rights lying to the north of the Trans-Canada Highway and south of Kamloops Lake in the event New Gold decides to sell these lands in the future separately and distinctly from the New Afton Project.

DECOMMISSIONING PLAN

Information on successful practices for reclamation from past mining operations at the New Afton site will assist in the design of future reclamation efforts. The New Afton mine decommissioning process will involve three distinct phases:

·	pre-decommissioning

·	decommissioning

·	post-decommissioning

The objective for all three phases will be to dispose of fixed machinery and buildings and to return the site to a condition acceptable to the Chief Inspector of Mines, MEMPR, as outlined in the Mines Act Permit. This work will be in accordance with the objectives of Part 10, Health, Safety and Reclamation Code for Mines in British Columbia. The Code requires that:

·	All machinery, equipment, and building superstructures shall be removed;

·	Concrete foundations shall be covered and revegetated unless, they have been exempted by the inspector because of demonstrated impracticality;

·	All scrap material shall be disposed of in a manner acceptable to the inspector.

PRE-DECOMMISSIONING PHASE

Decommissioning activities will commence in the final year of mine operations with establishment of the decommissioning management team and submission of a tailings decommissioning plan.

DECOMMISSIONING PHASE

The decommissioning process will be undertaken by a number of contractors with special skills or equipment. Those contractors are:

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·	A mechanical contractor to clean, dismantle and crate process and/or large pieces of mobile equipment;

·	An electrical contractor to disconnect, remove and crate for sale; controls, cabinets and transformers;

·	A civil contractor to dismantle buildings, prepare the tailings dam for shutdown, and remove-tailing lines;

·	A civil contractor to reclaim roads and waste dumps;

·	A scrap dealer to cut up structural steel structures that cannot be sold;

·	A crushing plant contractor equipped with a dozer/ripper, drop ball, magnet and crushing plant to crush the building floor slabs and piers;

·	A hydro seeding contractor to revegetate reclaimed areas;

·	An auctioneer to sell the process and mobile equipment, buildings and office furniture.

ENVIRONMENTAL MONITORING AND SURVEILLANCE

The stockpiles, Pothook pit earth dam, the visibility berm, and the dam slopes will be monitored for bare spots, slumping, and erosion. Any failures will be rectified. The stockpiles and the earth dam will be assessed on a monthly basis until the cover is stabilized and annually until it is required for reclamation. The surface cover of the berm will be assessed in the spring and fall for erosion. Any areas exhibiting erosion will be rectified.

The side slopes of the tailings storage facility will not be reclaimed until closure. However, slopes will be monitored on a monthly basis for water and wind erosion during operation. Surfaces exhibiting erosion will be amended.

Other reseeded areas such as the pipeline corridor and the access road side slopes will be monitored for surface erosion and reseeded where required.

POST-CLOSURE MONITORING

All reclaimed areas will require monitoring until vegetation is well established. Indicators of successful reclamation are high vegetative surface coverage and the absence of rills, gullies, and other erosional features. A soil specialist will be on-site periodically to provide guidance and make recommendations.

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The flat areas will be checked for surface coverage in the spring and fall for the first five years and annually for the next three years. Any issues related to plant establishment will be rectified. Records of inspection and a description of the conditions will be recorded on-site.

Reclamation of the TSF will require special attention because of the length of the slopes, the low fertility of the materials, the fine texture, the absence of soil structure, the potential for compaction, and the arid climate. Monitoring will be carried out regularly to ensure that reclamation has been successful. Areas exhibiting surface erosion will be amended. Records will be kept to assess the changes in percent bare ground or surface cover as well as plant health.

At closure, a range management plan will be developed for the areas which will be used for grazing. No animals will be allowed on the site for a minimum of three years to allow the bunch grasses to spread, establish, and produce seed. Range quality will be assessed in the spring and fall each year by a qualified professional who will determine range readiness. Range will be assessed every three years once it is in use to ensure the range plan is suitable.

All of the land which will be disturbed for the Project will be reclaimed. Traditional use in the area is cattle grazing on a landscape dominated by grasslands. The final end land use will be to develop grasslands which will be used for cattle grazing. Rangeland in British Columbia also includes wildlife use. Therefore, the final reclamation plan will include wildlife considerations, such as appropriate fencing and the development of shelter areas for various birds and small animals.

SUMMARY OF AREAS REQUIRING RECLAMATION

The stockpiles, Pothook Pit Dam, and the visual berm are classified as temporary infrastructures. Approximately 112.5 ha of flat surfaces will have received final reclamation treatment upon closure. As well, the TSF, which comprises 37 ha of surface and 4.6 ha of access road slopes, will be reclaimed. Therefore, a total surface area of 154.1 ha will have received final reclamation, with the remaining infrastructure treated to prevent wind and water erosion during the construction and operation of the mine.

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CLOSURE COST

For the 2007 Feasibility Study, Hatch made preliminary factored estimates of the cost of removing all equipment from the surface, demolishing the buildings, sealing the entrances to the underground, and regrading the site. The resale value of equipment and buildings was ignored. All underground equipment was left in place.

Rescan Environmental Services Ltd. (Rescan) estimated the cost of redistributing stockpiled till and top soil and the cost of revegetation. Rescan also assessed the cost of addressing remaining legacy liabilities from the previous operator and estimated the cost of post-closure monitoring (Rescan, 2007). All of the closure cost estimates were prepared in accordance with MEMPR guidelines and used the MEMPR spreadsheet. Table 19-12 contains the closure cost estimate.

TABLE 19-12 CLOSURE COST ESTIMATE

New Gold Inc. – New Afton Project

Demolition and Closure (Hatch)	C$ ‘000
Site Development and Roads	92
Process	4,268
Concentrate Transfer	17
Electrical Power	306
Tailings and Rock Disposal	648
Surface and Service Facilities	1,095
Mining	630
Subtotal	7,056

Reclamation (Rescan)	C$ ‘000
Tailings Dam	154
Road	2
Pothook Pit	4
Pothook Pit Earth Dam Footprint	0
Tailings Surface	3,432
Plant Site	299
Subtotal	3,891
Round-off	53

Total Cost	11,000

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TAXES

Scott Wilson RPA has relied on New Gold for guidance on applicable taxes, royalties, and other government levies or interests, applicable to revenue or income from the Project.

The taxes applicable to the New Afton Project are:

• BC net proceeds tax of 2%

• BC revenue tax of 13%

• BC income tax at a rate of 11%

• Federal income tax at a rate of 15%.

The BC net proceeds tax is taken against the revenue less operating costs including concentrate treatment and transportation. The agreed royalty to First Nations is not deductible.

The BC revenue tax is charged on gross proceeds less the cumulative expenditure account, which is the sum of operating and capital costs, with a 33% premium added to preproduction capital.

The federal and provincial income taxes are based on income less applicable capital allowances.

CAPITAL AND OPERATING COST ESTIMATES

The preproduction capital budget for the Project is summarized in Table 19-13. The capital is stated in Q2 2009 Canadian dollars. Sunk costs are not included in the capital summary. Capital cost estimates have been prepared by AMEC, New Gold, and AMC in the course of the detailed engineering work which was being done by AMEC until mid-2009 at which time the activity was essentially put on hold pending final financing. The cost estimate included sunk costs and Project costs, with over 5,300 line items with cost estimates for labour materials and contracts. Capital cost estimate revisions and adjustments continue as Project optimization tasks are still underway. The mine capital costs are built up from first principles for the mine development activities together with quotations for major equipment items.

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TABLE 19-13 CONSTRUCTION CAPITAL ESTIMATE

New Gold Inc. – New Afton Project

	C$ ‘000
Year	2010	2011	2012	Total
Site development	0	5,976	7,376	13,351
Process	2,542	55,035	31,678	89,255
Electrical	-	8,856	10,222	19,078
Tailings Disposal	-	6,903	3,063	9,966
Surface	-	9,414	1,890	11,304
Mining	37,474	148,837	48,179	234,491
Indirect	11,131	24,742	17,089	52,962
Contingency	-	24,416	6,104	30,520
Total	51,147	284,178	125,601	460,927

The capital cost details for the Project are shown in Tables 19-14 to 19-20.

TABLE 19-14 SITE DEVELOPMENT CAPITAL ESTIMATE

New Gold Inc. – New Afton Project

	C$ ‘000
Year	2010	2011	2012	Total
Access road	-	-	653	653
Portal Access Road (in pit)	-	-	-	-
Other Site Roads	-	-	1,476	1,476
Site Preparation & Drainage	-	-	5,246	5,246
Buried Services (excl electrical)	-	1,450	-	1,450
Fire Water System	-	1,411	-	1,411
Potable Water System	-	738	-	738
Fresh Water Supply	-	2,377	-	2,377
Total	-	5,976	7,376	13,351

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TABLE 19-15 PROCESS CAPITAL ESTIMATE

New Gold Inc. – New Afton Project

	C$ ‘000
Year	2010	2011	2012	Total
Building	2,542	33,228	16,614	52,384
Coarse Ore Reclaim	-	6,583	2,194	8,778
Grinding & Gravity	-	5,765	6,108	11,873
Flotation	-	3,337	2,224	5,561
Regrind	-	504	336	840
Concentrate Filtration	-	1,674	1,116	2,790
Tailings Prep & Pumping	-	2,422	1,615	4,036
Reagent Mixing	-	-	777	777
Plant Water System	-	659	439	1,098
Plant Air System	-	300	200	499
Truck Weigh Scale	-	564	-	564
Truck Wheel Wash	-	-	55	55
Total	2,542	55,035	31,678	89,255

TABLE 19-16 ELECTRICAL CAPITAL ESTIMATE

New Gold Inc. – New Afton Project

	C$ ‘000
Year	2010	2011	2012	Total
Main Substation	-	729	1,459	2,188
Standby Power Plant	-	-	70	70
Surface Power Distribution	-	2,628	1,314	3,942
Surface Communications	-	1,269	1,269	2,539
Main Underground Substation	-	-	1,880	1,880
Underground Power Distribution	-	4,229	4,229	8,459
Total	-	8,856	10,222	19,078

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TABLE 19-17 TAILINGS CAPITAL ESTIMATE

New Gold Inc. – New Afton Project

	C$ ‘000
Year	2010	2011	2012	Total
Tailings Earthworks	-	-	653	653
Tailings Delivery System	-	4,691	-	4,691
Water Reclaim System	-	2,211	2,211	4,423
Wells & Pumps	-	-	199	199
Total	-	6,903	3,063	9,966

TABLE 19-18 SURFACE SERVICES CAPITAL ESTIMATE

New Gold Inc. – New Afton Project

	C$ ‘000
Year	2010	2011	2012	Total
Dry and Offices	-	1,270	635	1,905
Workshop	-	2,605	-	2,605
Laboratory	-	1,901	951	2,852
Sewage Plant	-	466	-	466
Surface Mobile Equipment	-	2,983	-	2,983
Fuel Store	-	190	190	379
Security	-	-	115	115
Dry and Offices	-	1,270	635	1,905
Total	-	9,414	1,890	11,304

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TABLE 19-19 MINING CAPITAL ESTIMATE

New Gold Inc. – New Afton Project

	C$ ‘000
Year	2010	2011	2012	Total
Mine Access	4,561	4,652	1,339	10,552
Undercut	423	11,823	5,287	17,533
Extraction Level	991	15,290	11,219	27,500
Ore Transfer	-	4,963	2,855	7,818
Crushing	2,397	10,614	1,853	14,864
Conveying	7,817	20,923	5,641	34,380
Hoisting Facilities	2,582	6,754	2,368	11,704
Ventilation	7,206	23,623	4,602	35,430
Pit Dewatering	-	2,013	1,007	3,020
UG Pumping & Drainage	389	2,598	1,350	4,337
UG Services	1,528	4,773	1,284	7,584
Mining Mobile Equipment	-	28,475	-	28,475
Electrical Comms & Control	-	-	1,372	1,372
Geology & Diamond Drilling	-	173	165	338
Underground Road Construction & Maintenance	-	923	1,701	2,624
Mine Management & Tech Services	9,495	15,240	7,138	31,873
Adjustment	86	(4,000)	-1,000	(4,914)
Total	37,474	148,837	48,179	234,491

TABLE 19-20 INDIRECT CAPITAL ESTIMATE

New Gold Inc. – New Afton Project

	C$ ‘000
Year	2010	2011	2012	Total
Engineering & Procurement	1,753	2,004	751	4,507
Construction Management	-	4,164	2,082	6,247
Construction Indirects	-	4,995	2,497	7,492
First Fills	-	-	730	730
Spare Parts	-	1,313	657	1,970
Duties	-	240	120	360
Freight	-	2,657	1,328	3,985
Commissioning Costs	-	-	2,149	2,149
Owners costs	9,378	9,369	6,774	25,522
Contingency	-	24,416	6,104	30,520
Total	11,131	49,158	23,193	83,483

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The following is excluded from the capital cost estimate:

·	Project financing and interest charges

·	Escalation during construction

·	Any additional civil, concrete work due to the adverse soil condition and location

·	Insurance during construction

Taxes

·	Import duties and custom fees

·	Cost of geotechnical investigation

·	Working capital

·	Sunk costs to December 31, 2009

·	Exploration drilling

·	Costs of fluctuations in currency exchanges

·	Future expansion

·	Construction camp

·	Catering costs

The sustaining capital estimate for the Project is shown in Table 19-21.

TABLE 19-21 SUSTAINING CAPITAL BY AREA
New Gold Inc. – New Afton Project

		C$ ‘000
Mine Access		2,526
Undercut		28,431
Extraction Level		59,457
Ore Transfer		2,658
Hoisting Facilities		1,052
Ventilation		5,501
Diamond Drilling		4,051
UG Roads & Maintenance		7,637
Equipment and infrastructure		43,532
EPCM		631
Surface Mobile Equipment		4,777
Teck Royalty		12,000
Total Sustaining Capital		172,253

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TABLE 19-22 ANNUAL SUSTAINING CAPITAL

New Gold Inc. – New Afton Project

Year	C$ ‘000
2012	39,970
2013	35,881
2014	15,745
2015	24,372
2016	6,057
2017	6,490
2018	3,765
2019	15,256
2020	20,344
2021	3,575
2022	737
2023	62
Total	172,253

TABLE 19-23 ANNUAL SUSTAINING CAPITAL BREAKDOWN
New Gold Inc. – New Afton Project

	Access	Under- cut	Extrac- tion	Xfer	Hoist	Venti- lation	DD	Ser-vices	Equip-ment	EPCM	Surface	Royalty
2012	6	5,541	9,987	1,420	127	783	162	1,555	20,386	2
2013	8	9,932	18,158	804	167	1,066	321	2,451	1,038	5	1,933
2014	-	3,783	8,954	236	15	139	321	1,396	299	601	-
2015	-	997	6,047	-	-	-	996	681	2,347	20	1,284	12,000
2016	-	-	707	-	-	-	321	79	4,949	1	-
2017	-	-	-	-	-	-	321	-	6,083	1	85
2018	332	-	9	-	15	-	321	-	3,039	1	48
2019	2,181	3,645	2,704	197	493	3,449	321	-	1,709	-	557
2020	-	4,533	12,501	-	235	64	321	1,078	933	-	679
2021	-	-	392	-	-	-	321	397	2,380	-	85
2022	-	-	-	-	-	-	321	-	367	-	48
2023	-	-	-	-	-	-	5	-	-	-	57
	2,526	28,431	59,457	2,658	1,052	5,501	4,051	7,637	43,532	631	4,777	12,000

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The property operating costs are shown in Table 19-24. The costs are in Canadian dollars per tonne of ore milled. The costs do not include concentrate shipping or treatment, nor do they include corporate overheads. Operating costs have been compiled by New Gold and AMC. The costs for all areas were built up from first principles with a high level of detail in the cost build-up.

TABLE 19-24 OPERATING COSTS

New Gold Inc. – New Afton Project

	C$/Tonne Milled
	2012	2013	2014	2015	2016	2017	2018	2019	2020	2021	2022	2023	2024	Total
Mine	6.43	7.56	7.15	6.91	6.93	7.12	6.84	6.78	7.09	8.30	8.28	8.25	8.07	7.35
Mill	8.24	7.58	7.59	7.53	7.53	7.53	7.53	7.49	7.49	7.52	7.52	7.52	8.67	7.59
G&A	2.27	2.27	2.32	2.34	2.34	2.34	2.34	2.32	2.32	2.34	2.34	2.34	2.37	2.33
Total Cost	16.94	17.41	17.06	16.78	16.80	16.98	16.71	16.59	16.91	18.16	18.14	18.12	19.10	17.27

The site manpower is shown in Table 19-25. Note that the manpower levels are shown at the end of specific quarters when significant changes in manpower are planned to occur. The levels in this table include the owner’s team and underground development personnel, but not the construction force that will be required for underground and surface construction of the major facilities.

TABLE 19-25 MANPOWER

New Gold Inc. – New Afton Project

	2009	2010	2011	2012	2013	2014	2015	2016	2017	2018	2019	2020	2021	2022	2023	2024
Underground	64	71	271	263	194	144	116	140	172	172	130	130	130	130	121	130
Technical	11	11	23	23	19	16	16	15	15	15	16	15	15	13	12	12
Processing	8	9	12	74	80	80	80	80	80	80	80	80	80	80	80	80
G&A	20	23	23	23	23	23	23	23	23	23	23	23	23	23	23	23
Total	103	114	329	383	316	263	235	258	290	290	249	248	248	246	236	245

Management indicates that mine labour productivity has been adjusted for periods prior to any significant changes in activity to account for training and recruiting.

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PROJECT EXECUTION PLAN

The current Project schedule is not in a finished form, this is an aspect of the Project planning which Scott Wilson RPA recommends be addressed to generate a comprehensive Project development schedule including all of the changes in the Project since the 2007 Feasibility Study. The change in the production rate will have changed the tailings requirements and the faster mining rate may have impacted on the timing of the replacement access for the exploration access drift. At this time, the mine development is clearly the critical path activity.

At the time of the site visit, mine development was underway, the mill building had been erected, and the grinding mills were mechanically installed. Clearing and development of the TSF had commenced and the dewatering of the Afton pit to the Pothook pit was underway.

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ECONOMIC ANALYSIS

A Cash Flow Projection has been generated from the Life of Mine production schedule and capital and operating cost estimates, and is summarized in Table 19-26. A summary of the key criteria is provided below.

ECONOMIC CRITERIA

REVENUE

·	4 Mpta mining from underground

·	Mill recovery by zone, as indicated by testwork, averaging 87.8%.

·	Exchange rate US$1.00 = C$0.85.

·	Metal prices as shown in Table 19-10.

·	Net Smelter Return includes smelting, refining, transport, and penalty treatment costs.

·	Revenue is recognized at the time of mill production.

COSTS

·	Pre-production period: 30 months (January 2010 to July 2012).

·	Mine life: 12 years.

·	Life of Mine production plan as summarized in Tables 19-5 to 19-7.

·	Pre-production capital C$460.9 million.

·	Average operating cost over the mine life is $17.27 per tonne milled.

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TABLE 19-26 AFTER-TAX CASH FLOW

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CASH FLOW ANALYSIS

Considering the Project on a stand-alone basis, the undiscounted after tax cash flow totals C$707 million over the mine life, and simple payback occurs after approximately 4.5 years.

The after tax NPV at a 5% discount rate is C$374 million, and the IRR is 19%.

SENSITIVITY ANALYSIS

Project risks can be identified in both economic and non-economic terms. Key economic risks were examined by running cash flow sensitivities:

·	Metal prices

·	Operating costs (Total Cash Cost)

·	Capital costs

The 5% NPV after tax sensitivity over the base case has been calculated for -20% to +20% variations. The sensitivities are shown in Figure 19-13 and Table 19-27. The Project is most sensitive to metal price and less sensitive to capital and operating costs, which are almost equally sensitive.

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FIGURE 19-13 SENSITIVITY ANALYSIS

TABLE 19-27 SENSITIVITY ANALYSES

New Gold Inc. – New Afton Project

Sensitivity to Metal Price
Cu US$/lb	NPV @ 0% millions	NPV @ 5% millions	NPV @ 8% millions
1.64	302	104	28
1.85	505	239	136
2.05	707	374	244
2.26	910	510	353
2.46	1,112	644	460

Sensitivity to Operating Cost
$/tonne	NPV @ 0% millions	NPV @ 5% millions	NPV @ 8% millions
13.81	825	452	306
15.54	766	413	275
17.27	707	374	244
18.99	648	336	213
20.72	591	298	184

Sensitivity to Capital Cost
$ Millions	NPV @ 0% millions	NPV @ 5% millions	NPV @ 8% millions
513	835	484	345
577	771	429	295
641	707	374	244
705	645	321	195
769	581	266	145

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20 INTERPRETATION AND CONCLUSIONS

Scott Wilson RPA has carried out an audit of the Mineral Resource and Mineral Reserve estimates for the New Afton Project, and makes the following conclusions:

·	The drilling has been conducted in a manner consistent with industry standards.

·	The core handling, logging and sampling have been carried out in a manner consistent with industry standards.

·	Sampling and assaying has been conducted in a manner appropriate for the deposit type and mineralization, and using industry-standard practices.

·	Assay QA/QC protocols are consistent with common industry standards, and no evidence of persistent or systematic errors has been found. However, the QA/QC has been inconsistently monitored and requires more timely review.

·	Density measurements are made in sufficient numbers and with an appropriate methodology.

·	The database is generally free of serious errors although several inconsistencies have been found. At the time of writing, a new database system was being installed, which will implement more rigorous validation procedures.

·	The Mineral Resource estimate has been carried out using generally reasonable assumptions and parameters, and accepted practices.

Geological interpretation has resulted in improvements to the understanding of several features of the deposit, particularly the faulting and the DMZ. However, interpretations of the mineralized zone boundaries have not been maintained, resulting in the implementation of a grade shell approach to constraining the grade interpolation. In Scott Wilson RPA’s opinion, it would be preferable to base the resource boundaries on lithological and mineralogical characteristics, if possible.

·	The updated Mineral Resource estimate is not materially different from the 2006 estimate used in the 2007 Feasibility Study, and should not result in a major change to the Mineral Reserves.

·	The assay database for deleterious elements has been markedly improved since the last estimate but still lags behind the economic elements in terms of numbers of samples and coverage. Improvements could still be made to the block grade estimates for deleterious components.

·	The Mineral Resources should be quoted at a CuEq cut-off rather than a Cu cut-off.

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·	The Mineral Reserve is estimated to be 47.4 million tonnes of Probable Mineral Reserves grading 0.95% Cu, 0.69 g/t Au, and 2.03 g/t Ag.

·	The Mineral Reserves are considered to be amenable to underground mining through the use of block caving at a production rate of 4 Mtpa of ore.

·	The preproduction capital cost is estimated to be C$460.9 million plus sustaining capital of $172.3 million and not including sunk expenditures.

·	The estimated LOM operating cost is $17.27 per tonne milled.

·	Considering the Project on a stand-alone basis, the undiscounted after-tax cash flow totals C$707 million over the mine life, and simple payback occurs after approximately 4.5 years after the commencement of production.

·	The after-tax NPV at a 5% discount rate is C$374 million, and the IRR is 19%.

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21 RECOMMENDATIONS

Scott Wilson RPA makes the following recommendations:

·	Based upon the assumptions and estimates within this study, the continued development of the Project is recommended.

·	Database validation protocols should be enhanced and rigorously observed.

·	Geological interpretation and wireframing should continue with focus on the ore zones, in order to improve the constraints to the grade interpolations. If possible, the grade shell should be replaced by wireframes based on geological characteristics.

·	The Mineral Resource estimate should be quoted at a 0.4% CuEq cut-off. Changing the cut-off does not materially affect the Mineral Reserves.

·	Analytical work on the deleterious elements should continue in order to allow improvements to be made to the block models for these components.

·	The review of the project execution plan which is currently underway should be continued and completed including incorporation of all of the modifications since completion of the 2007 Feasibility Study.

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SCOTT WILSON RPA WWW.SCOTTWILSON.COM

22 REFERENCES

Behre Dolbear & Company, Ltd. (2003): Mineral Resource Estimate for the Afton Copper/Gold Project, Kamloops, B.C., 161 pp., filed on SEDAR.

Behre Dolbear & Company, Ltd. (2004): Advanced Scoping Study for the Afton Copper/Gold Project, Kamloops, B.C. Updated 2004, 187 pp., prepared for DRC Resources Corporation, filed on SEDAR.

Behre Dolbear & Company, Ltd. (2004): Mineral Resource Estimate for the Afton Copper/Gold Project, Kamloops, B.C., 160 pp., filed on SEDAR.

Bieniawski, Z.T. (1976): Rock Mass Classification of Jointed Rock Masses. Exploration for Rock Engineering. Z.T. Bieniawski Ed. Balkema, Johannesburg, pp. 97-106.

Darney, R., Friesen, R., Giroux, G. (2005): Summary Report on the 2003 and 2004 Exploration Program and Mineral Resource Estimate on the Rainbow Property, 55 pp., filed on SEDAR.

Hatch Ltd. (2007): New Afton Project Detailed Feasibility Study.

Konst, R. (2006): New Afton Project 2005-2006 Drilling Program, Sample Preparation and Analytical Quality Control Report, 17 pp., internal report to New Gold Inc.

Laubschner, D.H. (1990): A Geomechanics classification system for the rating of rocks in mine design, J. Of South African Institute for Mining and Metallurgy, Vol. 90, No. 10, Oct., pp. 257-273.

New Gold Annual Report (2008): filed on SEDAR.

New Gold (2009): Various Press Releases as filed on SEDAR, and the company website.

SGS Minerals Services (2007): An Investigation into the Recovery of Copper and Gold from the Afton Deposit

Sim, R., and Davies, B. (Draft 2009): Draft internal report to New Gold, 19 pp.

Wallis, S., and Giroux, G., (2006): Technical Report on the New Afton Project, 44 pp., internal report to New Gold Inc.

AMC Consultants Pty Ltd. (2007) Afton Project Feasibility Study, Underground Mining Study

AMC Consultants Pty Ltd. (2009), New Afton Project, Mineral Reserve Estimate, New Gold Inc.

Rescan Environmental Services Ltd. (2007), Application for a Permit Approving The Mine Plan and Reclamation Program Pursuant to the Mines Act R.S.B.C. 1996, C.293 New Afton Copper Gold Mine, British Columbia, Canada.

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Spreadsheets and internal documents provided by New Afton technical staff.

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23 DATE AND SIGNATURE PAGE

This report titled “Technical Report on the New Afton Project, British Columbia, Canada” and dated December 31, 2009, was prepared and signed by the following authors:

	(Signed & Sealed)

Dated at Vancouver, BC	R. Dennis Bergen, P.Eng
December 31, 2009	Associate Principal Mining Engineer

	(Signed & Sealed)

Dated at Vancouver, BC	David W. Rennie, P.Eng.
December 31, 2009	Principal Geologist

	(Signed & Sealed)

Dated at Vancouver, BC	Kevin C. Scott, P.Eng.
December 31, 2009	Principal Metallurgist

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24 CERTIFICATE OF QUALIFIED PERSON

R. DENNIS BERGEN

I, Raymond Dennis Bergen, P.Eng., as an author of this report entitled “Technical Report on the New Afton Project, British Columbia, Canada”, prepared for New Gold Inc. and dated December 31, 2009, do hereby certify that:

1.	I am Associate Principal Mining Engineer engaged by Scott Wilson Roscoe Postle Associates Inc. of Suite 501, 55 University Ave Toronto, ON, M5J 2H7.

2.	I am a graduate of the University of British Columbia, Vancouver, B.C., Canada, in 1979 with a Bachelor of Applied Science degree in Mineral Engineering. I am a graduate of the British Columbia Institute Technology in Burnaby, B.C. Canada, in 1972 with a Diploma in Mining Technology.

3.	I am registered as a Professional Engineer in the Province of British Columbia (Reg.# 16064) and as a Licensee with the Association of professional Engineers, Geologists and Geophysicists of the Northwest Territories (Licence L1660). I have worked as an engineer for a total of 27 years since my graduation. My relevant experience for the purpose of the Technical Report is:
	Ÿ	Practice as a mining engineer, production superintendent, mine manager, Vice President of Operations and a consultant in the design, operation and review of mining operations.
	Ÿ	Review and report, as an employee and as a consultant, on numerous mining operations and projects around the world for due diligence and operational review related to project acquisition and technical report preparation, including:
		¡	Engineering and operating superintendent at the Con gold mine, a deep underground gold mine, Yellowknife, NWT, Canada
		¡	Contribute to the Independent Report on the Reopening of the Cantung Mine at Tungsten, NWT, Canada
		¡	General Manager in Charge of the Reopening of the Cantung Mine, NWT, Canada
		¡	Detailed diligence review of the ERG Tailings Recovery Project, Ontario, Canada
		¡	VP Operations in charge of the restart of the Golden Bear Mine, BC, Canada
	Ÿ	Mining engineer in underground gold and base metal mines.
	Ÿ	Consulting engineer working on project acquisition and project design.
	Ÿ	Mine Manager at three different mines with open pit and underground operations
	Ÿ	Vice President of Operations responsible for an operating gold mine, development of a feasibility study for a project in Costa Rica and project review and evaluation related to project acquisition for numerous projects around the world including the successful acquisition of mines in Mexico and Australia and an interest in a mine in Argentina.

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4.	I have read the definition of "qualified person" set out in National Instrument 43-101 ("NI43-101") and certify that by reason of my education, affiliation with a professional association (as defined in NI43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI43-101.

5.	I visited the New Afton Project on September 3 and 4, 2009, and on October 23, 2009.

6.	I am responsible for preparation of Items 5, 18 and 19, as well as parts of Items 1, 2, 3, 17 (Mineral Reserves), 20, and 21 of the Technical Report.

7.	I am independent of the Issuer applying the test set out in Section 1.4 of National Instrument 43-101.

8.	I have had no prior involvement with the property that is the subject of the Technical Report.

9.	I have read National Instrument 43-101, and the Technical Report has been prepared in compliance with National Instrument 43-101 and Form 43-101F1.

10.	To the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Dated this 31st day of December, 2009

(Signed & Sealed)

Raymond Dennis Bergen, P.Eng.

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SCOTT WILSON RPA WWW.SCOTTWILSON.COM

DAVID W. RENNIE

I, David W. Rennie, P.Eng., as an author of this report entitled “Technical Report on the New Afton Project, British Columbia, Canada”, prepared for New Gold Inc. and dated December 31, 2009, do hereby certify that:

1.	I am a Principal Geologist with Scott Wilson Roscoe Postle Associates Inc. My office address is Suite 388, 1130 West Pender Street, Vancouver, British Columbia, Canada V6E 4A4.

2.	I am a graduate of the University of British Columbia in 1979 with a Bachelor of Applied Science degree in Geological Engineering.

3.	I am registered as a Professional Engineer in the Province of British Columbia (Reg.# 13572). I have worked as a geological engineer for a total of 30 years since my graduation. My relevant experience for the purpose of the Technical Report is:
	Ÿ	Review and report as a consultant on numerous exploration and mining projects around the world for due diligence and regulatory requirements
	Ÿ	Consultant Geologist to a number of major international mining companies providing expertise in conventional and geostatistical resource estimation for properties in North and South Americas, and Africa.
	Ÿ	Chief Geologist and Chief Engineer at a gold-silver mine in southern B.C.
	Ÿ	Exploration geologist in charge of exploration work and claim staking with two mining companies in British Columbia.

4.	I have read the definition of "qualified person" set out in National Instrument 43-101 ("NI43-101") and certify that by reason of my education, affiliation with a professional association (as defined in NI43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI43-101.

5.	I visited the New Afton Project on April 19 and 20, 2009 and again on September 3 and 4, 2009.

6.	I am responsible for preparation of Items 4 and 6 to 15 inclusive, as well as parts of Items 1, 17 (Mineral Resources), 20, and 21 of the Technical Report.

7.	I am independent of the Issuer applying the test set out in Section 1.4 of National Instrument 43-101.

8.	I have had no prior involvement with the property that is the subject of the Technical Report.

9.	I have read National Instrument 43-101, and the Technical Report has been prepared in compliance with National Instrument 43-101 and Form 43-101F1.

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10.	To the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Dated this 31st day of December, 2009

(Signed & Sealed)

David W. Rennie, P. Eng.

New Gold Inc. – New Afton Project	Page 24-4
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SCOTT WILSON RPA WWW.SCOTTWILSON.COM

KEVIN C. SCOTT

I, Kevin C. Scott P.Eng., as an author of this report entitled “Technical Report on the New Afton Project, British Columbia, Canada”, prepared for New Gold Inc. and dated December 31, 2009, do hereby certify that:

1.	I am a Principal Metallurgist with Scott Wilson Roscoe Postle Associates Inc. of Suite 388, 1130 West Pender Street, Vancouver, British Columbia, Canada V6E 4A4.

2.	I am a graduate of University of British Columbia, Vancouver, Canada in 1989 with a Bachelor of Applied Science degree in Metals and Materials Engineering.

3.	I am registered as a Professional Engineer in the Province of British Colombia (Licence # 25314) and the Province of Ontario (License # 90443342). I have worked as a metallurgical engineer for a total of 18 years since my graduation. My relevant experience for the purpose of the Technical Report is:
	Ÿ	Reviews and reports as a metallurgical consultant on a number of mining operations and projects for due diligence and financial monitoring requirements
	Ÿ	Process engineer at three Canadian base metals mineral processing operations
	Ÿ	Senior metallurgical engineer working for three multi-national engineering and construction companies on feasibility studies and in engineering design of mineral processing plants in Canada and South America
	Ÿ	Senior process manager in charge of process design and engineering for a metallurgical processing plant in South America

4.	I have read the definition of "qualified person" set out in National Instrument 43-101 (NI43-101) and certify that by reason of my education, affiliation with a professional association (as defined in NI43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI43-101.

5.	I have not visited the New Afton Project.

6.	I am responsible for preparation of Item 16 and parts of Items 1, 19, 20, and 21 of the Technical Report.

7.	I am independent of the Issuer applying the test set out in Section 1.4 of National Instrument 43-101.

8	I have had no prior involvement with the property that is the subject of the Technical Report.

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9.	I have read National Instrument 43-101, and the Technical Report has been prepared in compliance with National Instrument 43-101 and Form 43-101F1.

10.	To the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Dated this 31st day of December, 2009

(Signed & Sealed)

Kevin C. Scott, P.Eng.

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25 APPENDIX 1

This section contains the results of statistical analyses carried out on the composited drill hole data.

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