EX-99.1 2 techreport.htm TECHNICAL REPORT

CC Filed by Filing Services Canada Inc. 403-717-3898

New Gold Inc. - New Afton
NI 43-101 Independent Technical Report

Table of Contents

1.	Cover			i
2.	Table of Contents			ii
3.	Summary			1
	3.1	Property Description (This section was written by John Shillabeer P.Eng, Hatch)		1
	3.2	Ownership (This section was written by J. Shillabeer, P.Eng., Hatch)		3
	3.3	Resources (This section was written by David Rennie, P.Eng., Scott Willson RPA)		3
	3.4	Mining and Mineral Reserves (This section was written by Mike Thomas, MAusIMM (CP),
		AMC Consultants Pty Ltd)		3
	3.5	Metallurgy and Processing (This section was written by Ken Major, P.Eng., Hatch)		4
	3.6	Permitting (This section was written by Rolf Schmitt, P.Geo., Rescan Environmental
		Services Ltd.)		5
	3.7	Environmental (This section was written by Rolf Schmitt, P.Geo., Rescan Environmental
		Services Ltd.)		5
	3.8	Construction (This section was written by John Shillabeer, P.Eng., Hatch)		6
	3.9	Project Economics (This section was written by John Shillabeer, P.Eng., Hatch)		7
4.	Introduction ( This section was written by John Shillabeer P.Eng, Hatch)			10
5.	Reliance (This section was written by John Shillabeer P.Eng, Hatch)			13
	5.1	Property Tenures		13
	5.2	Marketing		13
	5.3	Taxes		13
	5.4	Other Geotechnical Engineering		13
	5.5	Pumping Condition Assessment		13
	5.6	Other		14
6.	Property Description and Location			15
	6.1	Introduction (This section was written by John Shillabeer, P.Eng., Hatch)		15
	6.2	Ownership		15
		6.2.1	Mineral (This section was written by NGD)	15
		6.2.2	Surface Ownership (This section was written by independent legal counsel)	17

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New Gold Inc. - New Afton
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7.	Accessibility, Climate, Local Resources, Infrastructure, and Physiography (This section was
	written by Rolf Schmitt, P.Geo., Rescan Environmental Services Ltd.)		20
	7.1	Accessibility	20
	7.2	Climate	20
	7.3	Local Resources and Infrastructure	20
	7.4	Physiography	21
8.	History (This section was written by David Rennie, P.Eng., Scott Wilson RPA)		22
9.	Geological Setting (This section was written by David Rennie, P.Eng, Scott Wilson RPA)		23
10.	Deposit Type (This section was written by David Rennie, P.Eng., Scott Wilson RPA)		27
11.	Mineralization (This section was written by David Rennie, P.Eng., Scott Wilson RPA)		31
12.	Exploration (This section was written by David Rennie, P.Eng., Scott Wilson RPA)		32
13.	Drilling (This section was written by David Rennie, P.Eng. of Scott Wilson RPA)		33
	13.1	2000-2003 Drill Programs	33
	13.2	2005-2006 Drill Programs	33
14.	Sampling Method and Approach (This section was written by David Rennie, P.Eng., of Scott
	Wilson RPA)		35
	14.1	2000-2003 Drill Programs	35
	14.2	2005 Underground Sampling	35
	14.3	2005-2006 Drill Programs	35
15.	Sample Preparation, Analyses, and Security (This section was written by David Rennie, P.Eng.,
	Scott Wilson RPA)		36
	15.1	2000-2003 Drill Programs	36
	15.2	2005-2006 Drill Programs	36
16.	Data Verification (This section was written by David Rennie, P. Eng., Scott Wilson RPA)		37
	16.1	2000-2003 Drill Programs	37
	16.2	2005-2006 Drill Programs	37
	16.3	Other (This section was written by John Shillabeer P Eng, Hatch)	38
17.	Adjacent Properties (This section was written by David Rennie, P.Eng., Scott Wilson RPA)		39

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New Gold Inc. - New Afton
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18.	Mineral Processing and Metallurgical Testing (This section was written by Ken Major, P.Eng.,
	Hatch)			40
	18.1	Introduction		40
	18.2	Program Description		40
		18.2.1	Sample Description	41
		18.2.2	Mineralogy	42
		18.2.3	Grinding	43
		18.2.4	Flotation	44
		18.2.5	Metallurgical Predictions	48
	18.3	Process Description		51
19.	Mineral Resource Estimation (David Rennie, P. Eng., of Scott Wilson RPA is responsible for
	sections 19.1 to 19.9 inclusive. Mike Thomas MAusIMM (CP), of AMC Consultants is
	responsible for section 19.10.)			58
	19.1	Introduction		58
	19.2	Wireframe Models		58
	19.3	Sample Database		59
	19.4	Capping of High Grades		59
	19.5	Compositing		61
	19.6	Bulk Density		62
	19.7	Geostatistics		62
	19.8	Block Models		63
		19.8.1	Search and Kriging Parameters	63
		19.8.2	Cut-Off Grades	64
		19.8.3	Block Model Results	64
		19.8.4	Block Model Validation	65
		19.8.5	Dilution Halo	68
		19.8.6	Arsenic and Mercury Model	70
		19.8.7	Classification	73
	19.9	Statement of Mineral Resources		74
	19.10	Mineral Reserves		75
		19.10.1	Sources of Information	75
		19.10.2	Preparation of the Resource Model for Reserve Estimation	75
		19.10.3	Initial Selection of Mining Outlines	77
		19.10.4	Estimating Block Cave Production	79
		19.10.5	Adjustment for Ore Losses	83
		19.10.6	Development Ore	84
		19.10.7	Total Ore Mined	85
		19.10.8	Mineral Reserve Classification	85
		19.10.9	Mineral Reserve Estimate	85

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20.	Other Relevant Data and Information (This section was written by John Shillabeer, P.Eng.,
	Hatch)			86
	20.1	Execution Plan		86
	20.2	Other information		88
21.	Interpretations and Conclusions			92
	21.1	General (This section was written by John Shillabeer P.Eng., Hatch)		92
	21.2	Project Economics (This section was written by John Shillabeer P.Eng., Hatch)		92
	21.3	Geology (This section was written by David Rennie, P. Eng., Scott Wilson RPA)		92
	21.4	Mining (This section was written by Mike Thomas, MAusIMM (CP)., AMC Consultants Pty
		Ltd.)		93
	21.5	Mineral Processing (This section was written by Ken Major, P. Eng., Hatch)		93
	21.6	Environmental Permitting (This section was written by Rolf Schmitt, P.Geo, Rescan
		Environmental Services)		93
22.	Recommendations			95
	22.1	Geology (This section was written by David Rennie, P. Eng., Scott Wilson RPA)		95
	22.2	Mining (This section was written by Mike Thomas, MAusIMM (CP)., AMC Consultants Pty
		Ltd.)		95
	22.3	Metallurgy and Processing (This section was written by Ken Major, P. Eng., Hatch)		96
	22.4	Surface Geotechnical (This section was written by Monte Christie, P.E., Vector Engineering
		Inc.)		97
	22.5	Environmental Permitting (This section was written by Rolf Schmitt, P.Geo., Rescan
		Environmental Services Ltd.)		97
	22.6	Groundwater (This section was written by Andrew Holmes P Eng., Piteau Associates)		98
	22.7	Construction (This section was written by John Shillabeer P Eng., Hatch)		98
	22.8	Project Opportunities (This section was written by John Shillabeer P Eng., Hatch)		98
		22.8.1	“C“ Zone Resource	98
		22.8.2	Early Ramp up to 4 Mt/y	98
		22.8.3	Delete Temporary Ore Crusher	98
		22.8.4	Optimize Cave Draw Schedule	99
		22.8.5	Two Tier Electricity Tariff	99
		22.8.6	Delay Pit Debris Stabilization	99
		22.8.7	Automation of Underground Operations	99
		22.8.8	Sub Level Caving Under The Block Cave	100
23.	References (This Section was Written by John Shillabeer, P.Eng., Hatch)			101

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24.	Date and Signature			107
	24.1	John Shillabeer, P.Eng., Hatch Ltd		107
	24.2	Ken Major, P.Eng., Hatch Ltd		108
	24.3	David Rennie, P.Eng., Scott Wilson RPA		109
	24.4	Andrew Holmes, P.Eng., Piteau Associates Engineering Ltd		110
	24.5	Monte Christie, P.E., Vector Engineering Inc		111
	24.6	Rolf Schmitt, P.Geo., Rescan Environmental Services Ltd		112
	24.7	Mike Thomas, MAusIMM (CP), AMC Consultants Pty. Ltd		113
	24.8	Mike Struthers, C. Eng, MAusIMM, MIMMM, AMC Consultants (UK) Ltd		114
25.	Additional Requirements for Technical Reports on Development Properties and Production
	Properties			115
	25.1	Mining Operations		115
		25.1.1	Mining Method	115
		25.1.2	Geotechnical Domains	117
		25.1.3	Structural Features	118
		25.1.4	Geotechnical data	118
		25.1.5	Cavability	121
		25.1.6	Fragmentation	124
		25.1.7	Subsidence	124
		25.1.8	Stability and Ground Support	126
		25.1.9	Mudrush Potential	126
		25.1.10	Pit Dewatering and Debris Stabilization	126
		25.1.11	Mine Layout and Access	130
		25.1.12	Block Cave Design	135
		25.1.13	Mine Production Rate	137
		25.1.14	Ore and Waste Transport	138
		25.1.15	Mine Ventilation	138
		25.1.16	Mine Infrastructure	139
		25.1.17	Emergency Egress	140
		25.1.18	Production Schedule	140
	25.2	Recoverability (This section was written by John Shillabeer P.Eng., Hatch)		143
	25.3	Markets and Transportation (H.M. Hamilton and Associates Inc.)		143
		25.3.1	Markets	143
		25.3.2	Transportation	144
	25.4	Environmental Considerations (This section was written by Rolf Schmitt, P.Geo., Rescan
		Environmental Services)		145
		25.4.1	Bond	145
		25.4.2	Remediation and Reclamation	145

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25.5	Taxes (This section was written by NGD and reviewd by PricewaterhouseCoopers LLP)		148
	25.5.1	Royalties	148
	25.5.2	Income Taxes	148
	25.5.3	Large Corporation Capital Tax	149
	25.5.4	Federal and BC Income Tax, and BC Mining Tax Rates	149
	25.5.5	BC Mining Taxes	149
	25.5.6	Flow-Through Funding Eligibility	149
25.6	Capital Cost (This section was written by John Shillabeer, P.Eng., Hatch)		149
	25.6.1	Introduction	149
	25.6.2	Estimate Summary	150
	25.6.3	Basis of Estimate	152
	25.6.4	Working Capital	160
	25.6.5	Sustaining Capital	160
	25.6.6	Closure Cost	160
	25.6.7	Contingency	161
25.7	Operating Cost (This section was written by John Shillabeer, P.Eng., Hatch)		162
	25.7.1	Summary of Mine Life Operating Costs	162
	25.7.2	Basis of Estimates	164
	25.7.3	Assumptions	165
	25.7.4	Exclusions	166
	25.7.5	Inclusions	166
	25.7.6	Mining (This section was written by AMC)	166
	25.7.7	Mineral Processing	168
	25.7.8	General and Administration	168
25.8	Economic Analysis (This section was written by Hatch)		168
	25.8.1	Cash Flow Model	168
	25.8.2	Metal Prices and Exchange Rate	169
	25.8.3	Economic Results	170
26.                    Illustrations			172

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NI 43-101 Independent Technical Report

List of Tables

Table	3-1	:	Mineral Resource Estimate (at C$10/t Cut-off)	3
Table	3-2	:	Mineral Reserve Estimate1	3
Table	3-3	:	Summary of Project Capital Costs by WBS Area (C$ ‘000)	7
Table	3-4	:	New Afton Total Operating Unit Costs	8
Table	3-5	:	Key Financial Assumptions and Economic Results	8
Table	4-1	:	Independent Qualified Person and Area of Responsibility	11
Table	4-2	:	Other Experts and Area of Responsibility	12
Table	18-1	:	Grinding Index Summary	43
Table	18-2	:	Locked Cycle Floatation Test Summary for Mesogene & Hypogene	45
Table	18-3	:	Locked Cycle Tests – Final Concentrate Assays	46
Table	18-4	:	Variability Samples – Cleaner Flotation Reagent Summary	47
Table	18-5	:	Drill Core – Variability Samples – Flotation Test Summary	47
Table	19-1	:	Sample Statistics	60
Table	19-2	:	Cutting Values	60
Table	19-3	:	Composite Statistics	62
Table	19-4	:	Block Model Geometry	63
Table	19-5	:	Metal Price and Recovery Factors	64
Table	19-6	:	Block Model Results	65
Table	19-7	:	Cross-Validation Results	66
Table	19-8	:	Comparison of Block and Declustered Composite Grades	66
Table	19-9	:	Comparison of Kriged and ID2 Estimates	67
Table	19-10:		As-Hg Sample Statistics	71
Table	19-11:		As-Hg Sample Statistics – Wireframe	71
Table	19-12:		Mineral Resources	74
Table	19-13:		C Zone Inferred Resource At US$1.20 Cu, US$450 Au, and US$5.25 Ag	74
Table	19-14:		Metallurgical Recoveries and Concentrate Grade Used to Estimate Mineral Reserves	76
Table	19-15:		Metal Prices and Other Parameters Used to Estimate Mineral Reserve	76
Table	19-16:		Description of Fields in the Combined Resource Model	77
Table	19-17:		Key Input Parameters to PC-BC Models	80
Table	19-18:		Summary Results from PC-BC Modelling	81
Table	19-19:		Key D50 Fragmentation Inputs to Cave-Sim (Main Ore Types)	82
Table	19-20:		Key D50 Fragmentation Inputs to Cave-Sim for Other Rock Types	82
Table	19-21:		Summary Results* from Cave-Sim Modelling	83
Table	19-22:		Results from Cave-Sim Expected Case After Losses*	84
Table	19-23:		Summary of Ore Recovered from Drawpoints	84
Table	19-24:		Development Ore	84
Table	19-25:		Ore Recovered from Development and Drawpoints	85
Table	19-26:		Mineral Resources Contained within the Vertical Projection of the Cave Footprint	85
Table	19-27:		Mineral Reserve Estimate1	85
Table	25-1	:	Geotechnical Statistics by Domain	119
Table	25-2	:	Virgin Stress Measurements Results	120
Table	25-3	:	Summarised RMR90 by Domain	122
Table	25-4	:	Ore Production Schedule	141
Table	25-5	:	Detailed Terms - Copper Concentrates	144
Table	25-6	:	Summary of Infrastructures Requiring Reclamation	147
Table	25-7	:	Capital Cost Estimate Summary by Area (C$ ‘000)	151
Table	25-8	:	Capital Cost Estimate Summary by Commodity (C$ ‘000)	152
Table	25-9	:	Lateral Development Costs by WBS	154
Table	25-10:		Vertical Development Rates (excluding muck removal)	154
Table	25-11:		Drawbell Drilling and Blasting Costs (excluding labour)	154
Table	25-12:		Construction Labour Rates	157

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NI 43-101 Independent Technical Report

Table	25-13:	Working capital invested during the project (US$ 000's)	160
Table	25-14:	Sustaining capital (C$ ‘000)	160
Table	25-15:	Closure Cost Estimate	161
Table	25-16:	New Afton Total Operating Unit Costs	163
Table	25-17:	Commodity and utility costs used in operating cost estimates	165
Table	25-18:	Average Mine Life Unit Operating Costs Excluding Electric Power and Heating (C$/tonne
		mined) (AMC)	167
Table	25-19:	Total Mineral Process Unit Operating Cost(C$/tonne Milled)	168
Table	25-20:	Summary of G&A Costs (C$/t milled)	168
Table	25-21:	Key Financial Assumptions	169
Table	25-22:	Metal Prices and Exchange Rate (3 Year Trailing)	170
Table	25-23:	New Afton Project Summary of Economic Results	170

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New Gold Inc. - New Afton
NI 43-101 Independent Technical Report

List of Figures

Figure	3-1	:	New Afton Location	2
Figure	3-2	:	Block Cave Outline Relative to the Existing Open Pit (Looking North)	4
Figure	3-3	:	Sensitivity of Before Tax NPV @ 5% to Variations in Costs and Metal Prices	9
Figure	6-1	:	Map – Mineral Titles (in Vicinity of Lease)	16
Figure	6-2	:	Map – Private and Crown Lands in the New Afton Mine Site Area	19
Figure	9-1	:	Regional Geology	24
Figure	10-1	:	3D Views of Deposit Note: Green=hypogene, Yellow=mesogene, Blue=supergene	29
Figure	10-2	:	Mineralized Zones	30
Figure	13-1	:	Underground Drift and Drill Plan	34
Figure	18-1	:	Hypogene Cu Recovery vs. Head Grade	48
Figure	18-2	:	Mesogene Cu Recovery vs Head Grade	49
Figure	18-3	:	Supergene Flotation Copper Recovery vs Head Grade	49
Figure	18-4		Hypogene Au and Cu Concentration Ratio	50
Figure	18-5	:	Mesogene Au Recovery vs Cu Recovery	50
Figure	18-6	:	Supergene Au Recovery vs Cu Recovery	51
Figure	18-7	:	Crushing and Grinding Simplified Flowsheet	53
Figure	18-8	:	Flotation and Dewatering Simplified Flowsheet	54
Figure	19-2	:	Tonnage Curve	65
Figure	19-3	:	Tonnage Curves For ID2 VS OK	67
Figure	19-4	:	Plan View of Preliminary Resource Block Model	78
Figure	19-5	:	Block Cave Outline Relative to the Existing Open Pit (Looking North)	79
Figure	20-1	:	Summary Project Development Schedule	89
Figure	25-1	:	Block Cave Outline Relative to the Existing Open Pit (Looking North)	117
Figure	25-2	:	Distribution of RMR76 by Domain (All Data)	120
Figure	25-3	:	Caving Stability Chart (after Laubscher, 1990)	123
Figure	25-4	:	Subsidence Limits Based on Empirical Assessment Methods	125
Figure	25-5	:	General View of Slope Conditions at Afton Mine Pit	127
Figure	25-6	:	Composite Plan of Mine Workings	131
Figure	25-7	:	Mine Plan Overlayed by Arial Photograph	131
Figure	25-8	:	Plan of B1 and B2 Undercut on 5085 Level	132
Figure	25-9	:	Plan of B1 and B2 Extraction Level on 5070 Level	132
Figure	25-10:		Plan of Ore Transfer Level and Crusher on 5055 Level	133
Figure	25-11:		Ventilation and Drainage Level on 5040 Level	133
Figure	25-12:		Block 3 Undercut Level on 4065 Level	134
Figure	25-13:		Block 3 Extraction Level on 4950 Level	134
Figure	25-14		Block 3 Ventilation and Drainage Drive on 4930 Level	135
Figure	25-15:		Undercut Design	136
Figure	25-16:		Plan View of Extraction Level and Drawpoint Design	137
Figure	25-17:		Scheduled Tonnage from Each Block by Quarter	142
Figure	25-18:		Scheduled Copper and Gold Grades by Quarter	142
Figure	25-19:		Scheduled Arsenic Grades by Quarter	143
Figure	25-20:		Life of Mine Unit Operating Cost	163
Figure	25-21:		Distribution of Mine Operating Costs Excluding Power And Heating (AMC)	167
Figure	25-22:		Sensitivity Of IRR Before Tax To Variations In Costs And Metal Prices	170
Figure	25-23:		Sensitivity of Before Tax NPV @ 5% to Variations in Costs and Metal Prices	171
Figure	26-1	:	Overall Site Plan	173
Figure	26-2	:	Surface Facilities – Mill Site Plan	174

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New Gold Inc. - New Afton
NI 43-101 Independent Technical Report

3. Summary

This technical report summarizes the results of the recently completed Feasibility Study of the New Afton Project for New Gold Inc. (NGD or the Company).

3.1 Property Description (This section was written by John Shillabeer P.Eng, Hatch)

The New Afton project is located approximately 10 km from Kamloops in South-Central British Columbia, Canada, (Figure 3-1). It is adjacent to the Trans Canada Highway, the Coquihalla Highway and 138 kV transmission lines owned by BC Hydro. Utility-owned natural gas and oil pipelines cross the property.

In 1999, DRC Resources Corporation (the Company’s former name), acquired mineral claims covering the former Afton open pit mine site. During 2000-2003, NGD mapped and sampled the open pit and drilled 90 diamond drill holes from the surface for a total of 42,450 m. In November 2004, NGD collared an adit in the south wall of the Afton open pit, drove 2,200 m of ramp and cross cuts, to delineate the ore zone and obtain metallurgical samples. NGD commissioned Hatch in December 2005 to coordinate the preparation of the detailed feasibility study. By April 2006, the cut-off date for the resource estimate in the Main zone, NGD had completed 78 underground drill holes for a total of 30,778 m.

NGD has obtained a mining lease covering an area of 902.9 ha and has signed a letter of intent with Teck Cominco Limited (Teck) to acquire the surface rights to the project lands and a water pipeline, from Kamloops Lake to the Project. The Company filed its application for a permit under the Mines Act in January 2007.

Since 2006, NGD has also pursued an active program of engaging with local First Nations, other stakeholders and government.

Summary
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New Gold Inc. - New Afton
NI 43-101 Independent Technical Report

Figure 3-1: New Afton Location

Summary
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New Gold Inc. - New Afton
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3.2 Ownership (This section was written by J. Shillabeer, P.Eng., Hatch)

The Company was granted a mining lease on November 29, 2006, denoted as Tenure No. 546063, and covering 902.9 hectares (the Lease). The Lease grants to the Company a 30-year renewable right to mine all Crown minerals, for an annual rental fee of C$9,750. In the area surrounding the Lease, there are ten 2-post legacy mineral claims, and 15 cell tenures covering 2,498 hectares. In addition, the Company has a further 3,304 hectares of claims to the north of the lease and 5,134 hectares of claims to the southeast.

The surface tenures and negotiations to acquire them are described in Section 6.

3.3 Resources (This section was written by David Rennie, P.Eng., Scott Willson RPA)

The Mineral Resource estimate published in September 2006 is summarized in Table 3-1.

Classification of the Mineral Resource estimate was carried out using the definitions in the CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines. Mineral Resources are assigned to one of three categories depending on the confidence level of the estimate

Table 3-1: Mineral Resource Estimate (at C$10/t Cut-off)

Measured	Tonnes	Cu	Au	Ag		DOLVAL
	Kt	(%)	(g/t)	(g/t)		C$/t
Measured	43,250	1.12	0.83	2.68	$	37.26
Indicated	22,410	0.84	0.66	2.42	$	28.34
Measured & Indicated	65,660	1.02	0.77	2.59	$	34.22

* Recovered value, assuming metallurgical recoveries of 90% for Cu and Au, and 75% for Ag, and a C$:US$ Exchange Rate of 0.88

The total Measured and Indicated Mineral Resources are 65.66 Mt at the C$10/t cut-off, with a total metal content of 954 mlbs Cu, 1,029 kozs Au, and 3,231 kozs Ag.

3.4 Mining and Mineral Reserves (This section was written by Mike Thomas, MAusIMM (CP), AMC Consultants Pty Ltd)

The Mineral Reserve estimate for the New Afton Project, published April 2, 2007 is presented in Table 3-2. The estimate has been prepared by AMC Consultants Pty Ltd (AMC) using the CIM standards on Mineral Resources and Mineral Reserves Definitions and Guidelines.

Table 3-2: Mineral Reserve Estimate¹

	Tonnes	Cu	Au	Ag	NSR  Value
	(Kt)	(%)	g/t	g/t	(C$/t)
Probable Ore Reserve	44.4	0.98	0.72	2.27	31.13

1 Estimated at US$1.45/lb Cu, US$475/oz Au using a cut-off NSR Value of $15/t of ore, recoveries for Cu varying between 80 – 93%; for Au varying between 69%-89% and a C$:US$ Exchange Rate of 0.88.

The reserve has been estimated using the resource block model prepared by Scott Wilson RPA, the contents of which form the basis of the Mineral Resource Estimate summarised in Table 3-1.

Summary
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New Gold Inc. - New Afton
NI 43-101 Independent Technical Report

The reserve has been estimated for a block caving mining method and takes account of the effect of mixing Measured and Indicated Resources with dilution from low-grade and barren material originating from within the cave outline and from overlying material. The Reserve also takes account of mineralized material that will be uneconomic to recover at the metal prices used for the reserve estimation and will remain in the cave at the end of the mine life.

Three cave areas are proposed. Block 1 (B1) and Block 2 (B2), separated by a low-grade pillar, will have drawpoints on the same extraction level (approximately 550m below surface), whilst Block 3 (B3) will have its extraction level approximately 120m deeper (Figure 3-2).

Ore from the extraction levels will be crushed in an underground crusher before being transferred to surface using a series of inclined conveyors installed in a ramp system. The ramp system will provide the main access to the mine and have a nominal gradient of 1:6 (17%).

A staged production build up is proposed, initially to an annualised rate of 1.6 Mtpa (4,400 tpd) for a nominal two-year period, increasing to a maximum production rate of 4.0 Mtpa (11,000 tpd). Mining will commence at the western end of B2 and progress to the east into B1. B3 will be mined in conjunction with B1

Figure 3-2: Block Cave Outline Relative to the Existing Open Pit (Looking North)

3.5 Metallurgy and Processing (This section was written by Ken Major, P.Eng., Hatch)

The mineral processing flowsheet that has been developed for the New Afton project has been based on the results obtained from the metallurgical test programs that were completed on exploration drill core samples and on metallurgical samples obtained from the underground development. The unit operations employed in the New Afton Project flowsheets are typical of other copper operations located in British Columbia.

Summary
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The metallurgical test program focused on the development of the process flowsheets using the metallurgical samples that New Gold recovered from the exposed underground ore faces for Mesogene and Hypogene ores. Ore variability tests were then completed using drill core samples representing special locations in the deposits and on blended samples, including blends of ore types and blends of ore and waste to represent the dilution effects associated with block cave mining. The results are discussed in Section 18.

The NGD mill was designed to process a blend of underground hypogene, mesogene and supergene ores. The process will utilize conventional techniques in progressive particle size reductions, physical separation and concentration of minerals from gangue into concentrates at marketable grades. Particle size reductions will be achieved with crushing, and two (2) stages of primary grinding, while mineral separations will be achieved by gravity concentration and differential flotation. A regrinding stage will be included in the flotation circuit. Processing will not involve chemical processes or alteration of the products. The low concentrations of reagents added during processing will largely be absorbed on particle surfaces to facilitate the physical separation and will not generate chemical reactions.

3.6 Permitting (This section was written by Rolf Schmitt, P.Geo., Rescan Environmental Services Ltd.)

The New Afton Project Application for a Permit Approving the Mine Plan and Reclamation Program under the B.C. Mines Act, was accepted for government technical review and public comment in January 2007. A number of additional government permits are required and will be expedited by the agencies through the lead role of the Chief Inspector of Mines under the Ministry of Energy, Mines and Petroleum Resources.

3.7 Environmental (This section was written by Rolf Schmitt, P.Geo., Rescan Environmental Services Ltd.)

The New Afton Project lies within the Interior Plateau and represents a heavily glaciated landscape overlying a region of complex geology in a semi-arid climate. The baseline environmental studies considered the brownfield nature of the site and the remnant, less-altered natural features. Two primary landscape types are:

a)	areas extensively disturbed (reconfigured and/or reclaimed) as a result of past mining activities and infrastructure development, and
b)	undisturbed landforms similar to those that existed prior to the development of the Afton Mine, and consist of Ponderosa Pine-bunchgrass vegetation interspersed by discontinuous alkaline waterbodies and ephemeral drainage altered by grazing land use.

Environmental studies conforming to provincial standards included: climate studies, air quality, meteorology, noise, emissions, surficial geology, soils, terrain, surface water quality and quantity, groundwater quality and quantity, hydrological modeling, pit lake bathymetry, aquatic biology, wetlands, terrestrial ecosystems, vegetation, invasive plants, wildlife including species at risk, land status and use, and land capability for agriculture, archaeology, metal leaching and acid rock drainage, impact assessments and development of environmental management plans.

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Water conservation is paramount in the dry interior grasslands of the Thompson Plateau. The project design will ensure that there is no surface water discharge from the site, and that all groundwater within the mining footprint will flow towards and collect in Afton Pit. The waste rock from the underground operations will be minimal and will be deposited in the bottom of the existing Afton Pit prior to flooding at closure. Nearly all waste rock has been determined as non-acid generating, therefore the final deposition of waste rock in the Afton pit, and subsequent flooding will ensure that the potential for any acid generation is highly unlikely. Tailings have also been determined to be non-acid generating and will be deposited in the southeast of the mining lease where all surface and groundwater will be directed to Afton pit, and the surface of the tailings facility will be revegetated at closure to support grazing and wildlife use.

3.8 Construction (This section was written by John Shillabeer, P.Eng., Hatch)

The project execution plan and schedule require approximately twenty-six months from project approval to achieve commercial production. Assuming the project is approved in May 2007 and construction begins forthwith, commercial production should begin in July 2009. The project schedule is governed by the time required for underground development. NGD has already taken steps to protect the schedule by selecting the mining contractor (Cementation Canada) and by ordering key components of the SAG mill which are scheduled for delivery in Q4 ’08. Initially the mine and process plant will operate for approximately two years at 4,400 t/d. The second part of the project will be implemented with an expansion between July 2010 and June 2011,enabling production to ramp up to 11,000 tpd by late 2011.

Cementation will be managed by NGD through an open book contract with risk sharing provisions. NGD intends that Cementation will be responsible for underground development, construction and mine production until the ramping up to 11,000 t/d is completed. An experienced EPCM contractor will be appointed and made responsible to NGD for the initial and expansion phases of the surface plant and infrastructure. NGD personnel will operate the surface plant and infrastructure throughout. Qualifications, assumptions and exclusions relevant to the capital cost estimate, operating cost estimate and economic analysis are set out in Sections 25.6, 25.7 and 25.8 respectively.

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3.9 Project Economics (This section was written by John Shillabeer, P.Eng., Hatch)

The estimated Project capital costs are summarized in Table 3-3 below.

Table 3-3: Summary of Project Capital Costs by WBS Area (C$ ‘000)

WBS	Description	Consultant	Project Phase		Total
			Initial	Expansion
A0	Site Devel./Roads	HATCH1	10,361	0	10,361
F0	Process	HATCH	56,300	20,749	77,049
G0	Conc. Transfer	HATCH	148	0	148
H0	Elec. Power	HATCH	5,550	0	5,550
J0	Tailings and Waste Disp.	HATCH2	5,721	9,385	15,106
K0	Surface Serv. Facilities	HATCH	10,594	3,320	13,914
M0	Mining	AMC3	125,902	78,673	204,574
		HATCH	15,536	3,668	19,204
P0	Capitalized Operating Cost	AMC	2,522	0	2,522
		HATCH	811	0	811
Z1	Engineering	AMC	4,891	2,430	7,321
		HATCH	12,041	2,949	14,990
Z2	Procurement	AMC	1,223	608	1,830
Z3	Const. Mgmt.	AMC	6,113	3,038	9,151
		HATCH	9,918	2,676	12,594
Z4	Const. Indirects	AMC	4,700	0	4,700
		HATCH	6,905	2,826	9,730
Z5	First Fills	HATCH	291	337	628
Z6	Spares	AMC	0	600	600
		HATCH	1,251	721	1972
Z7	Duties	HATCH	0	0	0
Z8	Freight	HATCH	2,694	1,137	3,831
Z9	Commissioning Costs	HATCH	876	688	1,564
Z10	Owner’s Costs Requiring Contingency	HATCH	6,261	1,071	7,332
	Total		290,607	134,875	425,482
	Contingency	HATCH	36,532	17,729	54,261
	Total		327,139	152,604	479,743
	Defined Owner’s Costs (no Contingency)		805	14,000	14,805
	Project Total Capital		327,944	166,604	494,548

1	Includes estimate by Urban Systems.
2	Includes Estimate by Vector
3	Includes estimate by MEG. (See Section 25.6 for explanation).

NGD will be required upon completion of signing the definitive agreement with Teck, to fund the purchase of the surface rights ($C16 million plus interest). NGD instructed Hatch to exclude the acquisition of the surface rights from project capital and the economic analyses. Working capital will be required in addition to the costs shown in this table. Initial working capital requirements are estimated to be C$9.6 million in 2009 and C$12.2 million in 2010.

Project capital costs were prepared in accordance with standard industry practice and the estimate is classified as a Class 3 estimate, as defined by the American Association of Cost Engineers, -5+15% accuracy. Sustaining capital costs and closure costs were estimated to be C$84.1 and C$11.0 million respectively. However, these costs are approximate and do not qualify as Class 3 estimates.

A summary of average operating costs is shown in Table 3-4.

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Table 3-4: New Afton Total Operating Unit Costs

	Initial	Expanded	Life of Mine
C$ Per tonne milled
Mining	11.80	4.61	5.14
Processing	7.98	4.02	4.33
G& A	2.03	0.99	1.08
Utilities	3.10	2.15	2.23
Total	24.90	11.76	12.76
US$ Per pound of payable copper
Mining	0.64	0.21	0.23
Processing	0.43	0.18	0.20
G & A	0.11	0.04	0.05
Utilities	0.17	0.10	0.10
Total	1.34	0.53	0.58

In addition, the cost of concentrate transport from the project site to a representative smelter in Japan is estimated to be US$99.12 per dry metric tonne.

Hatch carried out an after tax, discounted cash flow economic evaluation of the project in constant dollar terms as of January 1, 2007, assuming 100% equity financing, no leased equipment, and that concentrates will be transported to smelters in the far east via truck and rail through the port of Vancouver. The economic evaluation is based on the metals prices and exchange rate listed in the table. These are the averages for the last three years up to and including January 1, 2007. Hatch can offer no comment on the future of metal prices, exchange rates and cost inflation. These may have a significant impact on the project economics.

A summary of the results of the economic evaluation is shown below in Table 3-5 and Figure 3-3:

Table 3-5: Key Financial Assumptions and Economic Results

Copper price (LME)	US$/lb Cu	2.01
Gold price (LME)	US$/oz	487
Silver price (LME)	US$/oz	8.54
Exchange rate $C/$US		0.82
After tax cash flow	US$ million	396.4
NPV (net present value) @ 0%before tax	US$ million	614.3
NPV (net present value) @ 5% before tax	US$ million	265.9
NPV (net present value) @ 5% after tax	US$ million	143.0
Equity IRR (internal rate of return before tax)	%	13.6
Equity IRR (internal rate of return after tax)	%	10.4
Cash cost (average life of mine)1	US$/lb Cu	0.64
Project payback (from 2009)	years	6.3

Note ¹: net of by-product credits

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Figure 3-3: Sensitivity of Before Tax NPV @ 5% to Variations in Costs and Metal Prices

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4. Introduction ( This section was written by John Shillabeer P.Eng, Hatch)

NGD commissioned Hatch in December 2005 to coordinate the preparation of a detailed feasibility study for its New Afton Copper Gold Project (the Project). This report summarises the detailed feasibility study which was prepared to determine the technical feasibility, define the scope, estimate capital and operating costs and investigate the overall economics of developing the Project, as an underground mine and surface mineral processing facility. The Feasibility Study was to be sufficiently detailed to allow NGD to make a decision respecting the construction of the project.

NGD requested that Hatch prepare a Technical Report compliant with National Instrument 43-101 on the feasibility study based on the Project.

This report has been prepared using data obtained from diamond drilling, laboratory testwork, assays, vendor data and quotations, and data obtained from numerous prior reports, as detailed throughout the report.

This report is the product of technical contributions from the Consultants listed below. Consultants’ contributions are also identified in the text. Consultants were retained by NGD and reported to NGD. Hatch compiled all contributions to the feasibility report but did not supervise the preparation of, or verify, the information provided by other contributors to this report and takes no responsibility for any sections of this report that were prepared by persons other than Hatch. The following contributors to this report have conducted one or more visits to the site during the course of this work:

Monte Christie, PE	IQP, Tailings Design, Closure	Vector Engineering Inc.
Andrew Holmes, P.Eng	IQP, Hydrogeology	Piteau Associates Engineering Ltd
Ken Major, P.Eng	IQP, Metallurgy & Plant Design	Hatch
David Rennie, P.Eng	IQP, Resource Estimate	Scott Wilson Roscoe Postle
		Associates Inc
Rolf Schmitt, P.Geo	IQP, Permitting & Environmental Studies	Rescan Environmental Services
		Ltd
John Shillabeer, P.Eng	IQP, Study Manager	Hatch
Mike Struthers, C.Eng, MAusIMM,	IQP, Mine Geomechanical Engineering	AMC Consultants (UK) Ltd
MIMMM
Mike Thomas, - MAusIMM (CP)	IQP, Mine Engineering And Mineral	AMC Consultants Pty Ltd
	Reserve Estimate

The Qualified Persons taking responsibility for certain sections of this Report, and the extent of their responsibility for each section, for the purposes of National Instrument 43-101 are set out in the Table 4.1.

Also, for the purpose of NI 43 101, the Other Experts, and their areas of responsibility are listed in Table 4.2

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Table 4-1: Independent Qualified Person and Area of Responsibility

Responsible Person	Company	Primary Areas of Responsibility	Relevant Sections
Monte Christie, P.E.	Vector	Tailings Disposal: Methods Selection and	22.4, Information
	Engineering	Sequence, Geotechnical and Civil Engineering	relating to areas of
	Inc.	Design Integrated with Surface Water	responsibility in
		Investigations; Capital Estimate (civil	Sections 25.6 and
		quantities), Closure Planning.	25.7
		Plant Site Foundations: Site Investigation and
		Preliminary Design Advice
Andrew Holmes, P.Eng	Piteau	Hydrogeology: Site Investigations and	22.6, Information
	Associates	Assessments for Underground Mining and	relating to areas of
	Engineering	Tailings Disposal, Groundwater Flows	responsibility in
	Ltd	Modeling and Seepage Prediction	Sections 25.6 to
			25.7
Ken Major, P.Eng	Hatch Ltd	Metallurgical, Process Plant	3.5, 18, 21.5, 22.3
			Information
			relating to areas of
			responsibility in
			Sections 25.6 and
			25.7
David Rennie, P.Eng	Scott Wilson	Mining, Resource Estimates	3.3, 8 – 17(except
	Roscoe Postle		16.3), 19.1-19.9
	Associates Inc		inclusive, 21.3,
			22.1
Rolf Schmitt, P.Geo	Rescan	Environmental Investigations, Permit	3.6,3.7,7,21.6,22	.
	Environmental	Application Preparation,·Public Consultation	5,25.4,
	Services Ltd	Process Support	information
			relating to areas of
			responsibility in
			Sections 25.6 and
			25.7
John Shillabeer, P.Eng	Hatch Ltd	Study Compilation; Construction; Compiled	3.1,3.2,3.8,3.9,4,
		Capital and Operating Cost Estimates; Project	5,6.1,16.3,20,
		Economic Analyses.	21.1,21.2,22.7,22	.
		Supervised the Scope and engineering by other	8,23,Those
		engineers of underground mine: electrical	portions of 25.6,
		power distribution, water supply and	25.7 and 25.8 that
		dewatering, control and communications; also	are not expressly
		surface infrastructure civil, mechanical and	stated to be the
		electrical engineering.	responsibility of
			another party, 26.
Mike Struthers, C.Eng,	AMC	Geotechnical Database and Block Model,	25.1.2-25.1.9
MAusIMM, MIMMM	Consultants	Stresses, Location and Orientation of	inclusive
	(UK) Ltd	Infrastructure, Ground Support Design,
		Subsidence Prediction, Mud Rush Potential
		Analyses
Mike Thomas, MAusIMM (CP)	AMC	Mining Method, Equipment Selection,	3.4, 19.10, 21.4,
	Consultants	Development and Production Schedules,	22.2, 25.1.11-
	Pty Ltd	Reserve Estimate, Costs,.	25.1.18 inclusive,
			Information
			relating to areas of
			responsibility in
			Sections 25.6 and
			25.7

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Table 4-2: Other Experts and Area of Responsibility

Expert	Company	Area Of Responsibility	Relevant
			Section
	Independent legal counsel	Description Of Surface Tenure	6.2.2
Hugh Hamilton	H H Hamilton & Associates Inc	Concentrate Marketing And Transportation	25.3
Ender Parar P.Eng	Meg Consulting Ltd	Pit Debris Stabilization	25.1.10
Garry Eng	Price WaterhouseCoopers LLP	Tax Treatment In Economic Model	25.5
Peter Coxon P.Eng	Urban Systems Ltd	Condition Of Water Supply System	Included in
			25.6

DISCLAIMER

This report (the “Report”) was prepared by the qualified persons listed in Table 4.1 (the “QPs”). Each QP assumes responsibility for those sections or areas of the Report that are referenced opposite their name in Table 4.1. None of the QPs, however, accepts any responsibility or liability for the sections or areas of the Report that were prepared by other QPs.

The Report was prepared to allow New Gold Inc. (the “Owner”) to reach informed decisions respecting the development of the New Afton Project. Pursuant to its engagement with Hatch Ltd. the Owner is permitted to file the Report with the Canadian Securities Regulatory Authorities pursuant to provincial securities legislation. Except for the purposes legislated under provincial securities law, any use of the Report by any third party is at that party’s sole risk.

The Report is intended to be read as a whole, and sections should not be read or relied upon out of context. The Report contains the expression of the professional opinions of the QPs based upon information available at the time of preparation. The quality of the information, conclusions and estimates contained herein is consistent with the intended level of accuracy as set out in the Report, as well as the circumstances and constraints under which the Report was prepared, which are also set out herein.

As permitted by Item 5 of Form 43-101F1, the QPs have, in the preparation of the Report, relied upon certain reports, opinions and statements of certain experts. These reports, opinions and statements, the makers of each such report, opinion or statement and the extent of reliance is described in Section 5 of the Report. Each of the QPs hereby disclaims liability for such reports, opinions and statements to the extent that they have been relied upon in the preparation of the Report, as described in Section 5.

As permitted by Item 16 of Form 43-101F1, the QPs have, in the preparation of the Report relied upon certain data provided to the QPs by NGD and certain other parties. The relevant data and the extent of reliance upon such data is described in Section 16 of this Report.

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5. Reliance (This section was written by John Shillabeer P.Eng, Hatch)

In preparing its sections of this Report, Hatch have relied upon certain reports, opinions and statements of other experts. These reports, opinions and statements, the makers of each such report, opinion or statement and the extent of reliance is described below. Hatch hereby disclaims liability for such reports, opinions and statement to the extent that have been relied upon in the preparation of this Report, as described below. Hatch has relied on this information without independent verification.

5.1 Property Tenures

Section 6.2 describes the status of the land tenures in the Project area and the status of negotiations to acquire land for the Project. This section was written by independent legal counsel retained by NGD.

5.2 Marketing

H.M.Hamilton and Associates Inc. (Hamilton) were commissioned by NGD to complete the Concentrate Transportation and Marketing components of the feasibility study. Hamilton’s report sections advised the costs of truck, rail and ship transportation to Japan via Vancouver, including the cost of transhipment in Vancouver. Hamilton also provided advice on the future supply and demand for metals concentrates, typical smelter commercial terms, the importance of arsenic and mercury penalty elements, marketing strategy to address penalty elements, and net smelter return calculations. The economic model incorporates Hamilton’s estimated smelter terms and transport charges.

5.3 Taxes

Tax calculation routines in the economic model were prepared by NGD, PricewaterhouseCoopers LLP confirmed the validity of the calculations and the description in Section 25.5.

5.4 Other Geotechnical Engineering

BGC Engineering Inc were originally retained to provide geotechnical engineering services to the project. and were subsequently replaced by Vector Engineering Inc. (Vector). Vector had access to BGC’s key personnel and reports, but then conducted independent site investigations.

Marine and Earth Geosciences (MEG) Consulting Limited were retained by NGD to provide advice on the stabilization of the submerged debris at the bottom of the Afton open pit. Their preliminary engineering solution and report is summarized in Section 25.1.10. The engineering design will be reviewed following the pumping of the supernatant water out of the pit. The time required and approximate cost of stabilization were included by Hatch in the project schedule and capital cost estimates.

5.5 Pumping Condition Assessment

Urban Systems Ltd of Kamloops BC performed an inspection and provided a report on the condition of the water supply pump station and pipeline from Kamloops Lake to the property. Their report included a preliminary estimate of the cost to rehabilitate the facility and this estimate was included in the project capital cost estimate.

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5.6 Other

Hatch has relied on certain information supplied by NGD in the preparation of the capital cost estimate. This information (Owners costs) was estimated by NGD. It includes, for the duration of the project: the cost of the Owner’s management team, the rental and expenses associated with the Owner’s Kamloops office, financial management and accounting, construction and other insurance, and the cost of pumps to dewater the open pit. Hatch relied on this information without independent verification.

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6. Property Description and Location

6.1 Introduction (This section was written by John Shillabeer, P.Eng., Hatch)

In 1999, after Teck had relinquished their Afton claims and mining leases, NGD, (then called DRC Resources Corporation), acquired mineral claims covering the former Afton mine site and embarked on an exploration campaign. During 2000-2003, NGD mapped and sampled the open pit and drilled 90 surface diamond drill holes for a total of 42,450 m. In late 2003 NGD commissioned Behre Dolbear and Company to complete an advanced scoping study and resource estimate. The study outlined a potentially economically attractive project and concluded that advancing to the detailed feasibility study stage was warranted in conjunction with additional technical investigations. NGD collared an adit in November 2004 in the south wall of the Afton open pit and drove approximately 2,200 m of access ramp, drives and cross cuts to enable infill drilling of the ore zone, obtain samples for metallurgical testing and examine ground conditions. Results from the underground program were encouraging and as the outlook for copper and precious metal prices was positive, NGD commissioned a Feasibility Study in December 2005. By April 2006, NGD had completed 78 underground diamond drill holes, for a total of 30,778 m.

NGD has obtained a mining lease covering an area of 902.9 ha and has concluded a Letter of Intent with Teck to acquire the surface rights to the project lands and a water pipeline.

The New Afton deposit is a tabular body of porphyry-style Cu-Au mineralization that occupies a largely fault-bounded corridor, which traverses the property in an east-northeast/west-southwest direction. 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.

6.2 Ownership

6.2.1 Mineral (This section was written by NGD)

The Company has converted four post legacy claims (Afton 1, 2, 3, & 4) into a mining lease, which was granted on November 29, 2006 as tenure No. 546063 (the Lease). The Lease covering 902.9 hectares, grants to the Company the right for 30 years (renewable) to mine all Crown minerals. The annual rental under the Lease is C$9,750.

The block of claims surrounding the Lease comprises ten 2-post legacy mineral claims and 15 cell tenures covering an area of 2,498 hectares. In addition, the Company holds blocks of claims totalling 3,304 hectares to the northeast of this block and a block of claims totalling 5,135 hectares approximately 10 km to the southeast.

The Mining Lease and adjacent claims are illustrated on Figure 6-1.

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Figure 6-1: Map – Mineral Titles (in Vicinity of Lease)

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6.2.2 Surface Ownership (This section was written by independent legal counsel)

The New Afton Copper Gold Project consists of various parcels of land (collectively the “Project Land”) located within the Kamloops Land Title District (British Columbia), as more particularly described below and as approximately located and shown on the map (the “Map”) prepared by Rescan Environmental Services Ltd, dated February 27, 2007 and annexed hereto as Figure 6-2.

The following outlines the registered ownership interests in the Project Land and other surface rights in the Project Land:

(a)	those lands shown in brown on the Map and identified in the legend to the Map (the “Legend”) as Crown Land (the “Crown Lands”) are owned by Her Majesty the Queen in Right of the Province of British Columbia (the “Crown”);
(b)	those lands shown in yellow overlaid with horizontal purple lines on the Map are owned by Afton Operating Corporation (“AOC”) (the “Afton Lands”);
(c)	those lands shown in dark green overlaid with horizontal purple lines on the Map are owned by Teck Cominco Limited (“Teck”) (the “Teck Lands”),
	(the Afton Lands and the Teck Lands are together the “Fee Simple Lands”); and
(d)	those lands shown in two shades of green overlaid by diagonal dark green lines and identified in the Legend as “Grazing Lease 332483 26th April, 2007” and “Grazing Lease 333761 1st November, 2010” are owned by the Crown and are subject to the following grazing leases (together the “Grazing Leases”):
	(i)	those lands shown in green overlaid by diagonal dark green lines on the Map and identified in the Legend as “Grazing Lease 332483 26th April, 2007” are subject to Grazing Lease No. 332483 – File No. 0311961), made between the Crown, as lessor, and Sugarloaf Ranches Limited (“Sugarloaf”), as lessee, and for a term expiring on April 26, 2007; and
	(ii)	those lands shown in green overlaid by diagonal dark green lines on the Map and identified in the Legend as “Grazing Lease 333761 1st November, 2010” are subject to Grazing Lease No. 333761 – File No. 0095450, made between the Crown, as lessor, and Sugarloaf, as lessee, and for a term expiring on November 1, 2010.

Portions of the Project Land are subject to charges, liens and interests registered in the Kamloops Land Title Office that include, but are not limited to, covenants, easements and rights of way in favour of private and public corporations, Crown corporations, individuals and the Government of Canada (collectively the “Permitted Encumbrances”). Portions of the Fee Simple Lands are also subject to other registered interests including, but not limited to, mortgages in favour of Bank of Montreal, Royal Trust and other financial institutions (together, the “Financial Charges”).

The Trans-Canada Highway, shown in red on the Map, runs through a portion of the Project Land.

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Pursuant to a Letter of Intent dated January 8, 2007 (the “Teck LOI”):

(i)	Teck and AOC have agreed to sell to the Company the Fee Simple Lands; and
(ii)	Sugarloaf has agreed to relinquish to the Crown its interests in the Grazing Leases, as described in paragraph (d) above and as further identified in the Legend as “Grazing Lease Relinquished”.

The Teck LOI is conditional upon the Company, Teck, AOC and Sugarloaf executing a definitive agreement (the “Purchase Agreement”) that will concern items (i) and (ii) above, the discharge by Teck and AOC of the Financial Charges and obtaining all necessary regulatory approvals, all pursuant to the customary conditions for completing a transaction of this nature in British Columbia.

The Company has agreed to honour all pre-existing agreements made between Teck and third parties regarding access and rights of way over the Project Land and access to water, with the completion of the transactions contemplated by the Teck LOI all being subject to definitive agreements being executed to ensure these rights. The company also plans to acquire from Afton an easement, on similar terms as those set out in Easement No. T19493, as it relates to those lands known as District Lot 2017 upon which the water pipeline is located but which the Company is not purchasing from Afton.

The Company and Teck are in the process of finalizing the Purchase Agreement. To complete this acquisition, the Company will pay to Teck C$10 million upon closing, with an additional C$6 million to be paid (with applicable interest) any time within 2 years of closing. Teck will also be granted by the Company a 2% Net Smelter Return over the New Afton Copper-Gold Project, which the Company has the option to repurchase from Teck for C$12 million.

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Figure 6-2: Map – Private and Crown Lands in the New Afton Mine Site Area

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7. Accessibility, Climate, Local Resources, Infrastructure, and

Physiography (This section was written by Rolf Schmitt, P.Geo., Rescan Environmental Services Ltd.)

7.1 Accessibility

The New Afton Property is located on the south side of the Thompson River Valley, about 10 km west of Kamloops. The Trans-Canada Highway passes through the middle of the property, just west of its junction, with Highway 5 (the Coquihalla Highway). Access is by mine-site roads off the Trans-Canada Highway. NGD has entered into access agreements with Teck to cross that portion of their private land which lies within the NGD mineral claims.

Kamloops has an airport with daily air service to/from Vancouver and Calgary. Railroads belonging to both Canadian National Railway and Canadian Pacific Railway, service Kamloops with a line belonging to Canadian Pacific crossing the northern portion of the property near Kamloops Lake.

7.2 Climate

New Afton is located in the South-Central Interior of British Columbia which is characterized by warm summers where temperature 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 can drop to as low as –29°C.

The Kamloops area is in the rain shadow of the Coast Mountains and the climate is classified as semi-arid. 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. Evaporation rates are high and average 788 mm per year.

7.3 Local Resources and Infrastructure

Kamloops is a major transportation hub for highway, air and railroad facilities, and forestry, ranching, mining, and tourism are the most important economic activities in the area.

The proximity of Kamloops is of considerable importance to the New Afton Project. It is a natural resource-based city of 80,000 people with the Weyerhaeuser pulp mill and the nearby Highland Valley Copper Mine being significant local employers. There is positive support for mining activities.

The Project is served by infrastructure including: the Trans Canada Highway, high voltage electrical transmission line, water pipeline to Kamloops Lake, and a natural gas pipeline.

British Columbia Transmission Corporation (BCTC) plans, manages and operates B.C. Hydro’s electrical power transmission assets in the region, which include parallel 138 kV powerlines, IL 204 and IL 206 parallel to and just north of the Trans-Canada Highway crossing through the proposed New Gold mining lease area. Substations TMT (on IL 204) and KWD (on IL 206) are located at Kamloops. BCTC’s 2006 Capital Growth Plan envisions regional upgrades to the system to ensure stable load performance for future Project re-connection to the electricity grid.

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Infrastructure, and Physography
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Water for mining operations in excess of that currently available on-site will be pumped from Kamloops Lake. Pipeline infrastructure is in place to service the New Afton Project. NGD has arranged to purchase the water pipeline and pumphouse facilities from Teck as part of an Assets Purchase Agreement. NGD will be applying for a new water license to withdraw sufficient water for mining operations.

A natural gas pipeline owned and operated by Terasen Ltd. traverses the northern part of the Mining Lease and may provide a source of fuel for heating purposes.

7.4 Physiography

The Afton site has rolling topography with elevations ranging from 347 m above sea level (asl) at the water pipeline intake on Kamloops Lake, to a maximum of 790 meters on the Mining Lease, and 675 meters at the proposed mill site. The most significant features are the Afton and Pothook open pits, and the reclaimed waste rock dumps from the former Afton mine operations. The New Afton Project area was extensively glaciated resulting in deposits of glacial till and some lacustrine sediments of variable thickness on the property. Afton Pit, Pothook Pit and the man-made Waste Dump Pond contain an on-site water resource of about 2.9 million m³ water which will be conserved and used for early production.

The Thompson River (which widens into Kamloops Lake) is located along the northern periphery of the property.

Vegetation is typical of the semi-arid climate, consisting of grasslands, sagebrush and sparse groves of widely spaced pine trees.

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8. History (This section was written by David Rennie, P.Eng., Scott Wilson RPA)

Exploration in the Afton area began in the mid-1800’s, as prospectors pushed into the interior of British Columbia following the Fraser and Cariboo gold rushes. The Iron Mask property, staked in 1896, was the first in the Kamloops district. Mining was carried out from the turn of the 19^th century through to 1927 at a 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 1950’s, and 60’s.

Chester Millar acquired the property in the mid 1960’s, and formed a private company called Afton Mines Ltd. 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 New Afton deposit. For the next three years, over 150,000 ft of drilling was carried out by a number of operators. Duvall 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. 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 Mt grading 1% Cu, 0.016 oz/t Au, and 0.12 oz/t Ag (30.8 Mt grading 1% Cu, 0.58 g/t Au, 4.2 g/t Ag). Mining took place on the property at the Afton, Crescent, Pothook and Ajax pits. The mill closed in 1991, reopened again in 1994, and finally closing in 1997.

Seven deep diamond drill holes, drilled in 1978 and 1980 below, and to the southwest of the Afton pit, intersected 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.

In 1999, the Afton mining leases expired and the ground was staked by Westridge Enterprises Ltd., and Indo-Gold Development Ltd. NGD (then known as DRC Resources Corporation), 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, the company tabled a Scoping Study based on drilling results to that 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. An “Advanced Scoping Study” was completed in 2004.

In late 2004, a portal was collared on the south wall of the Afton Pit, and 2,200 m of decline and drift were driven to provide access for definition drilling and metallurgical sampling of the deposit. The drifting was completed by September of 2005. Diamond drilling, both from surface and underground was completed in December, 2006. Total exploration expenditures on the New Afton project by DRC and NGD to the end of 2006, were $C39.1 million.

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9. Geological Setting (This section was written by David Rennie, P.Eng, Scott Wilson RPA)

The New Afton deposit is hosted within the Cherry Creek member of the Iron Mask batholith complex (see Figure 9-1). 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 (see Figure 9-1). At New 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, with 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.

Nicola Group rocks border the Cherry Creek intrusive on the west and southwest of the deposit, but are not principal 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 cross-cut 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 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 (south) side of the picrite is limited.

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, at times, are difficult to discriminate from Cherry Creek rocks. 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 in the Kamloops strata are visible in the north wall of the New Afton Pit.

Geological Setting
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Figure 9-1: Regional Geology

Geological Setting
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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-ore processes. Zones of intersection between deep seated NW-SE-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 NW SE 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 Foot Wall (west) 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 New Afton pit, 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 much of the deposit, especially in the hypogene zone.

Subsequent fault movements re-shuffled portions of the deposit, in places 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 (ca. 400 m below surface in places).

Lastly, during the Tertiary, tensional stress developed in a NW-SE direction resulted in NE-SW-trending graben structures and the development of depositional basins for Kamloops Group sediments.

Important fault structures in the deposit area, as described by Brian Bower and Brian O’ Connor , NGD’s geologists, are listed below:

Primary Fault Structures

• Hanging Wall Fault – oriented 023⁰/78⁰ SE. Steeply-dipping strike-slip fault comprising seams of clay gouge and breccias. Clay gouge seams are typically less than a few metres thick, however, the fault is defined by abundant brecciation, carbonate veining and strong clay/carbonate/sericite alteration. This fault is locally well mineralized. When the Hanging Wall Fault is in contact with the Picrite unit, it is generally represented by variable massive clay gouge and sub-rounded fault breccias. Latite post mineral dykes are often found within and adjacent to the Hanging Wall Fault. These dykes are generally brecciated and sheared when caught up in the fault.

• Foot Wall Fault – oriented 023°/78°SE to 036°/74NW Generally parallels the Hanging Wall Fault, but consists primarily of carbonate and clay infill material adjacent to in-situ vein brecciation. Several drill holes have been lost in the attempt to drill across this structure.

Secondary Structures

• North Easters (000° to 015°) / (76° to 90°) SE a series of parasitic smaller fault structures that appear to be parallel to the Hanging Wall Fault.

• Picrite West (054°/81° SE) a series of splay faults along the Hanging Wall Fault.

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• North Westers (320°/76° SW) a series of northwest trending steep faults within the intrusive, sub-parallel to the Nicola-Intrusive contact (perhaps parallel to the Cherry Creek fault zone).

• Northern (075° to 090°) / 60° SSE a series of faults south of the Hanging Wall Fault, possibly offsetting the Foot Wall and Hanging Wall Faults’ transitional ore zones.

Structural orientations mapped from surface and underground were plotted on a stereonet and two dominant structural trends were identified. These dominant trends are planes measuring 086°/48°S and 130°/45°SW. This structural data includes both joints and faults combined.

The complex structural nature of the New Afton deposit is exemplified by the high degree of jointing and fracture filling seen in the underground and pit exposures. These joint sets and fracture fills are ubiquitous throughout the deposit area.

Geological Setting
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10. Deposit Type (This section was written by David Rennie, P.Eng., Scott Wilson RPA)

The New Afton deposits are Cu-Au silica-saturated, alkalic porphyry style deposits. Mineralization resulted from late stage hydrothermal activity driven by remnant heat within an alkalic intrusive complex. Thermal gradients within these systems gave 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 concentrated fluid flow.

At New Afton, 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. Cu-Au 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.

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 New Afton deposit is a tabular body of porphyry-style Cu-Au mineralization that occupies a largely fault-bounded corridor, which traverses the property in an east-northeast/west-southwest direction (see Figure 10-2). This corridor is bounded on either side by the Hanging Wall (SE) and Foot Wall (NW) Faults. The ore body does not come in contact with the Foot Wall 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.

The interpretation of the general geometry of the zones has not changed much with recent drilling. The bulk of the deposit occupies the Main Zone, which consists of a tabular mass measuring 900 m long by approximately 100 m wide and spanning a vertical distance of about 350 m (See Figures 10-1 and 10-2). The zone dips vertically to steeply south-southeast, and plunges at moderate angles to the southwest. Two subsidiary satellite bodies (Hanging Wall Zones), occur to the southeast 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 60⁰ SE).

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 the destruction of 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.

Deposit Type
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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.

The same fault zones controlled and localized weathering, which has resulted in widely varying degrees and distribution of supergene enrichment. NGD geologists have recognized mineralogical changes brought on by weathering and have interpreted the shape of the weathered zones. Primary hypogene sulphide mineralization occurs as disseminated and fracture-filling chalcopyrite and bornite. The exterior boundary of this zone is defined as the first appearance of Cu sulphide accompanied by disappearance of magnetite-pyrite. The hypogene zone comprises the westernmost half of the deposit.

Mesogene (or mixed) zone mineralization is defined by the first appearance of chalcocite. The chalcocite can occur as total replacements of, or rims surrounding chalcopyrite and bornite. The mesogene zone occupies a central “plug” within the Main Zone and encompasses both satellite bodies.

The appearance of native Cu defines the boundary of the supergene zone. The supergene zone occupies the north-eastern portion of the Main Zone. It is somewhat unconventional in distribution as it does not form a weathered cap over top of the primary sulphide mineralization. Fractures have conducted weathering fluids downwards in a relatively confined area but to significant depths (>500 m in places).

Deposit Type
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Figure 10-1: 3D Views of Deposit*
Note: Green=hypogene, Yellow=mesogene, Blue=supergene

* Top looking north, middle looking south and bottom looking east.

Deposit Type
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Figure 10-2: Mineralized Zones

Deposit Type
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11. Mineralization (This section was written by David Rennie, P.Eng., Scott Wilson RPA)

Economic mineralization within the hypogene zone comprises fine-grained disseminated chalcopyrite and relatively minor bornite. Mesogene zone mineralization consists of chalcopyrite and bornite that have been variably altered to chalcocite. The supergene zone contains secondary native Cu concentrated along fractures, along with remnant chalcocite. Sulphide minerals within the supergene zone have been largely weathered out.

Native Au is rarely observed. Au and Ag reportedly occur as electrum grains within the chalcopyrite and bornite.

Mineralization
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12. Exploration (This section was written by David Rennie, P.Eng., Scott Wilson RPA)

Prior to NGD’s involvement, the only work that had been done on or near the deep resource at Afton deposit was the seven holes drilled by Afton Mines Ltd. in 1978 (2 holes) and 1980 (5 holes). DRC mapped and sampled the pit (as well as any available outcrop surrounding the pit) and drilled 90 surface diamond core holes for a total of 42,450 meters (139,272 ft.) during the period 2000 to 2003.

Diamond drilling commenced from the adit in February 2005 and, to the cut-off date for the Main Zone Mineral Resources estimate (April 2006), the most recent drilling totalled 30,778 m in 78 holes. This brought the total amount of drilling on the project to 77,836 m in 171 holes. Drilling continued to the end of 2006.

Five exploration holes totalling 2,996 m were drilled in the Pothook Pit area. These holes are not associated with the main mineralized zones and do not impact the Mineral Resource estimate.

Exploration
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13. Drilling (This section was written by David Rennie, P.Eng. of Scott Wilson RPA)

13.1 2000-2003 Drill Programs

NGD used Atlas Drilling Company 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 (which produces 5 cm drill core), except for four drill holes which were completed with BQ rods (4 cm drill core).

From 2000 to April 2003, a total of 90 diamond drill holes were collared and the data from 82 drill holes has been used for resource estimation and geological mapping (Figure 13-1). 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 mineralized zone was estimated to be 365 m long, 76 m wide (average), and 300 m in vertical height.

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, averages 90 m wide, and 300 m in vertical height.

13.2 2005-2006 Drill Programs

The 2005 underground drilling program was carried out by Boisvenu Drilling of BC using NQ2 equipment, producing 5 cm diameter core. 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 24,864 m in 66 holes during 2005. Drilling continued into 2006, and as of April 2006, 12 underground holes totallying 5,194 m and a further 10 surface holes were drilled totalling 3,627 m. Additional drilling has continued to delineate the mineralization, to collect geotechnical data, and for condemnation purposes.

The core was brought from the drill 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 both geologically and geotechnically, including percent recovery, RQD, character and number of fractures. Geological observations include rock type, mineralization, and alteration. The boxes were photographed prior to being sampled. Core recovery was generally good.

Underground drilling has been completed on sections and consisted of several drill hole fans from each station. Based on the drilling to date, the main mineralized body appears to be about 1,000 m in length, 50 to 100 m wide and 400 m in height with a moderate plunge to the west.

Drilling
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Figure 13-1: Underground Drift and Drill Plan

Drilling
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14. Sampling Method and Approach (This section was written by

David Rennie, P.Eng., of Scott Wilson RPA)

14.1 2000-2003 Drill Programs

NGD 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, diamond sawed, and sampled with half the core retained in the core box at the Company’s core shack. Samples were conveyed by lab employees from site to Eco-Tech Laboratories Ltd. of Kamloops, B.C. for analysis for Cu, Au, Ag and Pd. Core from the 2000 program was split using a blade splitter rather than diamond sawed and was not photographed prior to splitting. From time to time, one-foot un-split mineralized sections were retained for structural studies.

Sample lengths were predominantly 2.0 or 3.05 m (10 ft), 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 were 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, all of which 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.

14.2 2005 Underground Sampling

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

14.3 2005-2006 Drill Programs

Each drill hole was sampled in 2 m lengths or less, as marked by the geologist. Scott Wilson RPA notes that, while the sampling protocol included provision for sampling at less than 2 m lengths, virtually all of the sampling done since the start of 2005 has been on 2 m intervals. Sampling was reportedly based on visible mineralization but in practice, most of the hole was sampled. The core was sawn in half, the samples bagged, then put into plastic pails for shipping. The sampler completed the sample tags and for every 20 samples, included a blank sample from previously drilled barren andesite. A tag was entered for a standard to be put into the sample stream at the lab, and in addition, one sample was ¼ split to provide a duplicate.

The pails were kept in a locked trailer 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 ½ core was stored in the original boxes on pallets outside. Each pallet was covered with plastic wrap and securely fastened.

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

Scott Wilson RPA is of the opinion that the sampling was carried out to industry standards.

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15. Sample Preparation, Analyses, and Security (This section was written by

David Rennie, P.Eng., Scott Wilson RPA)

15.1 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 sub- samples, and pulverized to 95% -140 mesh.

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

• Au was sub-sampled to 30 gram aliquots and analyzed with conventional fire assay using atomic absorption (AA) and/or Inductively Coupled Plasma (ICP) finish. Minimum reported detection for Au was 0.005 g/t.

• Cu was determined by AA using aqua regia digestion. “Metallic” Cu (when required) included 2 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 NGD in the 2000-2003 programs was assayed by Eco Tech Laboratories of Kamloops, BC. Eco Tech are Certified Assayers, participate in the National Canmet Proficiency Testing, and maintain their own in-house Quality Assurance and Quality Control (QA/QC) program. They have been in the analytical testing business for over 27 years, and are familiar with assaying the Afton samples.

The 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 NGD’s field office until picked up by personnel from Eco Tech Laboratories for transport to their facilities. Drill core is stored in core racks at the locked, secure core shack. Rejects are securely stored at Eco Tech’s office, and pulps are securely stored at NGD’s field office.

15.2 2005-2006 Drill Programs

Stewart Wallis of Scott Wilson RPA visited the Eco Tech Laboratory on June 7, 2005, inspected the facilities, and discussed the assaying methodology and QA/QC protocols with Jutta Jealouse, President. Internal checks consisted of a minimum 2 repeats, 1 blank, 2 re-splits, and 2 or 3 reference standards, one for Cu, one for Ag or one combined Cu/Ag and one for Au/Pd. If native Cu was reported on the sample sheets, a metallic screen analysis was run in addition to the regular assay.

The lab was clean and proper procedures were in place to track each sample. Assay results and internal check results were checked and batches rerun if problems observed. The laboratory was in the process of obtaining ISO certification.

Cu and Ag assays were determined using standard acid digestion followed by AA. Au and Pd were determined using fire assay followed by an AA finish.

As noted above, the samples were kept in a locked shed until picked up by the laboratory for transport. Scott Wilson RPA had no concerns regarding security or the integrity of the samples.

Sample Preparation, Analyses and
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16. Data Verification (This section was written by David Rennie, P. Eng., Scott Wilson RPA)

16.1 2000-2003 Drill Programs

NGD established a QA/QC program in order to ensure that assay lab results were within accepted industry standards:

• Assay standards were routinely used to control assay precision;

• One in nine pulp samples were re-assayed by Eco Tech;

• One in twenty-five 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.

16.2 2005-2006 Drill Programs

Scott Wilson RPA has not collected independent samples from the property as it has been previously visited and reported on by a number of independent Professional Geologists and is a former producing mine.

NGD has continued to maintain a QA/QC program consisting of the addition of standards, duplicates and blanks into the sample stream. The protocols are 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. A total of 18,508 unprepared samples were submitted to Eco Tech Laboratory 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 analysed by Ron Konst, P. Geo., an independent consultant retained by NGD. Mr. Konst noted that there were 53 blanks assays, and 18 standards assays with results outside of an acceptable error limit. These samples comprised 7% and 2% of the total blanks and standards 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 recommended that the batches with out-of-spec 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 out-of-spec blanks were found to be misidentified in the database and did not represent improper assay results.

The duplicate data were analysed 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.

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Scott Wilson RPA reviewed the Konst report and the QA/QC data, and conducted independent analyses of the QA/QC results. The duplicate analyses were subject to Thompson-Howarth plots to confirm the results of the Konst report. As well, each of the blanks assays were plotted against the sample that immediately preceded it in the sample stream to check for evidence of contamination in sample preparation. No evidence of sample contamination was indicated by this analysis. Scott Wilson RPA reviewed the plots of the standards assays and confirms that there was no indication of systematic bias. Review of the blanks and standards plotted in chronological order did not show any trend or particular time period where assays tended to be out-of spec.

Scott Wilson RPA concurs with the principal conclusion of the Konst report. In Scott Wilson RPA’s opinion, the assay data is suitable for use in Mineral Resource estimation.

16.3 Other (This section was written by John Shillabeer P Eng, Hatch)

Hatch has relied on certain information supplied by NGD in the preparation of the capital cost estimate. This information (Owners costs) was estimated by NGD. It includes, for the duration of the project: the cost of the Owner’s management team, the rental and expenses associated with the Owner’s Kamloops office, financial management and accounting, construction and other insurance, the cost of pumps to dewater the open pit. Hatch relied on this information without independent verification.

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17. Adjacent Properties (This section was written by David Rennie, P.Eng., Scott Wilson RPA)

The Project is the most advanced exploration and development project in the area. The ground adjacent to the southeast is held, and is being explored by Abacus Mining and Exploration Corp. As of June 2005, Abacus had drilled 43 diamond drill holes totalling 16,829 m on the property (Darney, et al, 2005). Porphyry-style Cu and Au mineralization has been intersected in these drill holes. Mineral Resources on the property are reported to be 101 Mt in the Indicated category grading 0.24% Cu and 0.07 g/t Au, with an additional 10.6 Mt of Inferred grading 0.18% Cu and 0.05 g/t Au

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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18. Mineral Processing and Metallurgical Testing (This section was written by

Ken Major, P.Eng., Hatch)

18.1 Introduction

A metallurgical test program was developed for NGD’s Project by Hatch in conjunction with New Gold’s technical representatives and SGS Lakefield. Hatch and NGD staff developed a detailed metallurgical sampling program and protocol, which was carried out by NGD staff. Hatch have no reason to believe that this sampling deviated from the agreed program and protocol. The results obtained by testing these samples were in the ranges expected, based on experience with similar deposits. The primary test work was completed by SGS-Mineral Services (SGS). Secondary programs were completed by Pocock and Knelson. Mineralogy was completed by Advanced Mineral Technology Laboratory (Amtel) and Vancouver Petrographics. This work was conducted from March to October 2006. 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. These ore types and their proportions within the resources are:

• Supergene [9%]

• Mesogene [39%]

• Hypogene [52%]

The preliminary metallurgical evaluations focused on the Mesogene and Hypogene ore types. The metals of economic interest were copper, gold, and silver. In the Supergene and Mesogene ores a minor amount of copper values occur as native copper [Cu]; but the majority of copper values occur as secondary sulphides in the form of Chalcocite [Cu2S] and Bornite [Cu5FeS]. In the Hypogene ore, the copper sulphide is primarily Chalcopyrite [CuFeS2]. Gold values were found in the free state, as electrum, associated with copper sulphides and Pyrite [FeS], as sulphide binaries and in the rock. The mode of occurrence for silver has not been determined.

18.2 Program Description

This program was conducted in two phases. The first phase was conducted on metallurgical samples representing the Mesogene and Hypogene ore types. The first phase established metallurgical requirements as they represented the bulk of the tonnage in the deposit. These materials were collected by NGD geological staff from an exploration decline into the deposit. A crosscut was developed from the main decline to intercept the Mesogene and Hypogene zones. The Metallurgical sample program was subsequently revised to include barren hanging and foot wall materials adjacent to the various ore types in order to monitor their dilution effects in the flowsheet.

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The second phase evaluated the variability of metallurgical response to standard conditions. This phase of the program employed split drill core samples from various spatial locations in the orebody. The Supergene ore type was also added to the variability program to determine the response to the selected flowsheet. This phase of the program included split drill core samples from diamond drill holes 90 and 91. Drill hole 90 represented type 1 and type 2, Supergene ore types. Drill hole 91 more completely represented the Mesogene ore. SGS completed grinding test work on the Mesogene and Hypogene ores. All of the bulk and variability flotation tests were completed by SGS.

Pocock Industrial Inc. of Salt Lake City, Utah was engaged to complete the product dewatering tests. They conducted their work on rougher flotation tailings samples of Hypogene, blends of Mesogene/Supergene and Mesogene/Mesogene hanging rock types. In their scope of work, they performed flocculant screening, static and dynamic gravity sedimentation, and pulp rheology, pressure and vacuum filtration tests. SGS Lakefield conducted solid-liquid settling tests on flotation concentrates.

Amtel completed a mineralogical evaluation and reported on the deportment of gold values in the three ore types. Vancouver Petrographics Ltd. issued two mineralogical examination reports on 30 metallurgical samples for the two main ore types and on 70 core and metallurgical samples representing the three ore types, foot wall and other rock units.

18.2.1 Sample Description

First Phase Metallurgical samples

First phase metallurgical samples were collected from a decline cross cut drift located below the former Afton Mine open pit. The underground bulk sample locations were selected by NGD’s geologist based on face sample assays. Hatch metallurgists viewed the proposed sample locations with the geologist. NGD coordinated the collection, labelling and shipping of the samples from the mine-site to SGS.

The metallurgical samples represent the two main ore types, the Mesogene (transition) and the Hypogene (primary). The ore faces were sampled to include cobble or fist sized chunks. For each ore type there was a division to represent low, medium and high grade ranges as follows:

• High Grade 2.5 to 3.5% copper equivalent;

• Medium Grade 1.5 to 2.0% copper equivalent; and

• Low Grade 1.0 to 1.5% copper equivalent

For each of the Mesogene and Hypogene metallurgical samples, nine 300 kg drums were collected and shipped to SGS Lakefield for metallurgical testing. For each ore type, there were three sample drums for each grade range.

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Dilution Metallurgical samples

Dilution samples were submitted for ore blending to represent the selected block caving mining method. Block caving will result in lower selectivity at the ore interfaces. To evaluate dilution effects on the metallurgical response barren hanging and foot wall samples were submitted to SGS-Mineral Services. A total of 15 plastic pails of rock and split drill core samples weighing approximately 33 kg per pail were collected for this work. The dilution type samples represented the Mesogene, Hypogene and Picrite rock types.

Second Phase Drill Core Samples

Second phase variability samples were selected to represent various spatial locations within the ore zone. For this work, 18 pails of split drill core samples representing Supergene, Mesogene and Hypogene and three grade ranges were submitted to SGS-Mineral Services.

Supergene Drill Core Samples

Supergene half split drill core samples were also included to represent the characteristics for this ore type. Six pails of samples from drill hole 90 were provided to SGS of which three pails were for Type 1 and three pails for Type 2 Supergene ores. Type 1 Supergene represents a stronger hematite alteration and Type 2 Supergene represents a weaker hematite alteration but with more interval fractures and stringer controls.

Drill Hole 91 Core Samples

Drill hole 91 core samples were also provided to SGS-Mineral Services. Nine pails of core sample from drill hole 91 represent Mesogene ore. These additional core samples were submitted to provide the expected range of primary and secondary copper minerals that should be present for the three ore types.

18.2.2 Mineralogy

Three organizations conducted mineralogical studies on NGD’s Project samples. SGS Mineral Services conducted a modal analysis on the Mesogene and Hypogene ores and a mineralogical study on Hypogene, third cleaner copper concentrate samples. This SGS study identified the mineral Tennantite [Cu11FeAs4S13] as the primary source of Arsenic. Amtel conducted a gold deportment study for the Supergene, Mesogene and Hypogene ores. Vancouver Petrographics Ltd. (VPL) conducted two mineralogical studies. One for Hypogene and Mesogene on the underground metallurgical samples and a petrographic study on 70 drill core and underground samples for all three ore types and foot wall rocks. The Vancouver Petrographics studies were briefly summarized as follows:

Hypogene ore - VPL identified chalcopyrite as the copper mineral of interest with a chalcopyrite : pyrite ratio of 25:1 with the chalcopyrite occurring primarily as free grains but with inclusions in silicates. Gold particles varied in size from 5 to 40 microns and occur as native gold between pyrite and chalcopyrite grains. There was also a sericite clay mineral which was confirmed in the SGS modal analysis.

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Mesogene ore, VPL found that chalcopyrite, bornite and chalcocite are the copper minerals of interest, in varying proportions that occur as intergrowths that replace the original chalcopyrite mineralization. It was noted that chalcocite tends to be finer grained and dispersed. The fine grained particles will have a negative influence on the final concentrate grade. In this ore type pyrite is extremely rare. The content of sericite is higher and the carbonate is non-reactive to hydrochloric acid. No gold particles were observed in these polished thin sections

Supergene ore, VPL noted the main copper minerals were chalcocite and native copper. Native copper occurred as microfracture coatings, which will render recovery by flotation difficult. Lower levels of sericite and carbonate compared to the Mesogene ore were observed.

Foot Wall Zone, VPL found these rocks resembled Mesogene rocks, but with lower levels of sericite. Some sulphides were present as barren pyrite. Copper minerals were rare.

The SGS-Mineral Services modal analysis quantified the levels of various minerals. The program indicated the presence of higher levels of Tetrahedrite [4Cu2S.Sb2S3] in the Mesogene rock type. The concentration of this mineral could cause higher Sb levels in the copper concentrate but the flotation test program did not identify any economic concerns. As noted above, arsenic also occured in these ores, primarily as Tennantite. Its level in Mesogene posed an economic concern for the concentrate produced from this ore type. The levels of sericite and chlorite were high enough in both the Mesogene and Hypogene rock types to cause slime problems in the dewatering of flotation tailings.

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.

18.2.3 Grinding

A detailed effort was devoted to determining the grinding characteristics for the Mesogene and Hypogene ore types. Grind studies were initiated on coarse - rock type and copper grade classified samples. In these coarse sizes, energy determinations were completed using the Bond Low-energy Impact Test and the JKTech Drop-weight Test [DWT]. The coarse rock type - grade samples were then stage crushed to 1-1/4” riffled and homogenized to form two grindability rock type [average grade] composites. The grindability composite was further stage crushed to six mesh, and at the appropriate stages samples were removed for the 18” Mill Autogenous work index [MacPherson test] (AWI), the Bond Rod Mill (RWI), the Bond Ball Mill work index (BWI) and Abrasion Index (AI) determinations. Those results have been summarized in Table 18-1

Table 18-1: Grinding Index Summary

	AWi	RWi	BWi	Ai
Rock Type	(kWh/t)	(kWh/t)	(kWh/t)	(g)
Mesogene	18.20	18.30	19.80	0.071
Hypogene	19.20	18.50	21.80	0.164

These Work Index values have classified both ores as being hard and competent. For the Abrasion index classifications, the Mesogene ore was considered to be mildly abrasive and the Hypogene ores as abrasive.

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18.2.4 Flotation

Metallurgical samples

The first phase test work on the bulk sample evaluated: primary and regrind product sizes; collector and frother types; flotation kinetics; and the open circuit and locked cycle metallurgy.

Included in this work was an evaluation of the dilution effects on the metallurgy by inclusion of barren country rock in the flotation feed. A limited amount of flotation work was conducted on the Supergene ore, because it only represented 5 to 10% of the mineable mineral ore reserve tonnage and will only influence the stages of mine and mill production. 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 will be returned for regrinding. The rougher tailings and the cleaner-scavenger tailings will be handled separately but combined for calculating the final tailings. Potassium amyl xanthate [PAX], Cytec’s 3418A [a phosphine reagent] collector; MIBC, DF250 and Pine Oil frothers, were the main flotation reagents evaluated in this study.

Dilution Tests

Dilution tests on metallurgical samples were conducted to determine the effect on metallurgy, when additions of barren country rock were included in the mill feed. The adoption of a block caving mining method initiated this work. A five step rougher kinetic series was conducted on Hypogene and Mesogene ore types, with the addition of ten and twenty percent hanging wall and foot wall materials. That was followed by a series of cleaner tests on Mesogene that was diluted by Supergene ore types 1 and 2, (between 20 and 40 percent). The results indicated that the diluting rock produced similar rougher copper and gold recovery, but at higher mass recoveries and lower concentrate grades. This would increase the load on the regrind mill and cleaner circuit.

Locked Cycle Tests

Locked Cycle Tests were conducted on metallurgical samples in duplicate for the main ore types. Within each duplicate set, a different frother reagent [MIBC or DF250] was used. The results below show that MIBC will provide a higher concentrate grade but a lower copper recovery than DF250.

The locked cycle test results are provided in Table 18-2.

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Table 18-2: Locked Cycle Floatation Test Summary for Mesogene & Hypogene

Test	Prim. & Regr.	Reagents	Combined	Weight	Assays		Distribution
No.	(K80µm)	(g/t)	Product	(%)	Cu %	Au g/t	Cu %	Au g/t
LCT-M1	~160	PAX(21)	Cu 3rd Cl Conc	4.12	28.6	17.8	88.6	79.8
		MIBC(67.5)	Cu 1st Cl Scav Tail	8.40	0.530	0.65	3.3	6.0
[Mesogene]			Cu Rougher Tail	87.82	0.152	0.152	10.1	14.5
[Cycles C-F)			Combined Final Tail	96.22	0.185	0.20	13.4	20.5
			Calc Head.	100.3	1.33	0.92	102.0	100.3
LCT-M2	~160	PAX(22.5)	Cu 3rd Cl Conc	4.83	25.2	16.2	89.4	84.4
		3418A (5)	Cu 1st Cl Scav Tail	10.80	0.350	0.39	2.8	4.5
[Mesogene]		DF250 (25)	Cu Rougher Tail	84.19	0.113	0.120	7.0	10.9
[Cycles C-F)			Combined Final Tail	95.00	0.139	0.15	9.8	15.4
			Calc Head.	99.8	1.36	0.93	99.2	99.8
LCT H1		PAX(10)	Cu 3rd Cl Conc	3.90	29.4	15.30	91.6	84.9
		MIBC(47.5)	Cu 1st Cl Scav Tail	5.47	0.39	0.39	1.72	3.04
[Hypogene]			Cu Rougher Tail	90.6	0.083	0.087	6.02	11.2
[Cycles C-F]			Combined Final Tail	96.1	0.10	0.10	7.74	14.3
			Calc Head.	100.0	1.25	0.71	99.4	99.2
LCT H2		PAX(10)	Cu 3rd Cl Conc	4.88	24.5	14.30	93.8	93
		3418A (5)	Cu 1st Cl Scav Tail	8.28	0.27	0.18	1.77	1.95
[Hypogene]		DF250 (47.5)	Cu Rougher Tail	86.0	0.063	0.043	4.26	5.0
[Cycles D-F]			Combined Final Tail	94.3	0.06	0.04	6.03	6.9
			Calc Head.	100.0	1.28	0.75	99.9	99.9

A summary of locked cycle – third cleaner concentrate assays have been provided in Table 18-3. Assays for Drill hole 91 locked cycle test have also been included for the record. There will be gold and silver credits; arsenic and antimony will be present in the concentrate. The insoluble values were low for the ore type locked cycle tests but they are high in the Drill hole 91 concentrates. Higher insoluble values would have been expected due to the high clay content of these ore types. Flotation difficulties due to insoluble materials were briefly mentioned in one bulk sample test.

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Table 18-3: Locked Cycle Tests – Final Concentrate Assays

From Locked Cycle Tests

Test		LCT-M1	LCT-H1	LCT-H91-1
Cycles		C-F	C-F	D-F
Rock Type		Mesogene	Hypogene	Mesogene
Metal	(%,g/t)
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
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	200.0	100.0	N/A
Insol	%	12.1	7.1	29.6

N/A = Not Available

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. Due to the block caving mining method, the development work has focused on the development of one flowsheet for all ore types.

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. The economic advantage for operating a gravity concentration circuit has not been identified from the test work. Based on successes at porphyry copper gold ores, a gravity concentration circuits will be included in the primary and regrind circuits.

Variability Samples

Variability work for the second phase of the flotation test program was conducted on drill core samples to determine the variability of flotation response on a spatial basis throughout the ore zones. Various intervals from sixteen drill holes were provided for this evaluation. Within these intervals the samples were classified by ore type and by high, average or low copper grades.

A summary of the variability test reagent requirements have been included in Table 18-4.

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Table 18-4: Variability Samples – Cleaner Flotation Reagent Summary

		Reagent Addition Ranges (g/t)
Rock Types	Collectors		Frothers		Modifiers
	PAX	3418A	DF250	Pine Oil	Na2S
Supergene	30 to 60		77.5 to 115		50
Mesogene	12.5 to 37.5	5 to 10	32.5 to 207.5	0 to 6
Hypogene	9.75 to 14.75	5	15 to 22.5	0 to 3
DDH 91	28.5 to 33.5	0 to 5	27.5 to 40	0 to 3

An overall summary of the variability test results has been presented in Table 18-5. All variability samples that were tested achieved acceptable copper concentrate grades.

Table 18-5: Drill Core – Variability Samples – Flotation Test Summary

	Calc Head		Rougher Flotation
Test Number	Cu	Au	Distribution (%)
	(%)	(g/t)	Cu	Au
VS-F	2.47	0.94	93.2	91.2
VS-A	1.79	1.38	86.3	87.6
VS-B	1.79	0.50	91.0	87.0
VS-D	1.64	0.40	92.1	82.3
VS-C-G1	1.38	0.16	86.5	67.2
VS-C	1.26	0.17	86.2	69.8
VS-E-R	0.96	0.35	91.5	83.2
VS-E	0.87	0.34	89.2	76.5
VM-F	2.78	0.64	90.5	90.3
VM-C	2.10	0.71	93.1	91.5
VM-D	1.57	0.36	89.7	82.0
VM-A	1.12	1.06	93.2	90.3
VM-E	0.78	0.18	84.1	78.2
VM-B	0.63	0.65	84.9	86.7
VH-D	2.06	1.31	93.0	90.3
VH-B	1.90	1.15	89.8	92.7
VH-B-R	1.86	1.21	90.9	90.9
VH-E	1.35	1.06	94.2	95.1
VH-A	1.25	1.04	94.5	93.3
VH-F-R	0.60	0.50	91.3	88.5
VH-F	0.58	0.55	91.1	86.7
VH-C	0.55	0.59	90.9	94.0
V-H91-C	1.74	1.44	92.9	91.5
F56(Comp)	1.25	1.21	91.1	85.8
V-H91-B	1.14	1.61	88.9	84.2
V-H91-A	0.94	1.20	93.2	93.5

VS=Variability Supergene VM=Variability Mesogene

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18.2.5 Metallurgical Predictions

The data from the open cleaner variability tests have been analysed to develop equations for projecting the recovery of copper, gold and silver and the deportment of the penalty elements, arsenic and mercury, in the concentrate for all 3 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 3 ore types. Recoveries have been normalized on a 28% Cu concentrate grade for Hypogene which contains primarily chalcopyrite, on a 30% Cu concentrate grade for Mesogene because of the occurrence of bornite, and 58% Cu concentrate for Supergene due to the presence of native copper.

In all cases a higher offset of 2% to 3% has been applied to the best fit equations based on batch open circuit data, to project plant performance. This offset was supported by locked cycle test results for Hypogene and Mesogene. Typically, locked cycle recovery will be higher than that from batch open cleaners for the same concentrate grade and it will indicate the recovery that may be achieved in an optimized plant operation. For Supergene, the correlation has been developed based on flotation only and the offset will allow for potential native copper recovery by gravity concentration.

For Hypogene, the correlation for copper recovery in a 28% Cu concentrate has been shown in Figure 18-1. This correlation may be applied to head grades of 0.5% Cu to 1.5% Cu, the range of head grade tested. Further test work will be required to determine if the correlation applies to head grades outside this range.

It was noted that the recovery in two tests for around 2% Cu head grade in Figure 18-2 were lower than expected. The lower recoveries were attributed to the coarser grinds that were observed in these tests.

Figure 18-1: Hypogene Cu Recovery vs. Head Grade

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For Mesogene, a correlation based on tails losses for a 30% copper concentrate provided the best recovery representation as shown in Figure 18-2. The correlation may be applied to the tested head grades of 0.5% Cu to 2% Cu. The data for a low-grade sample has been excluded from the analysis because it contained 23% native copper which was not recovered by flotation. As in the case with Hypogene the correlation has been off set from the batch test data, based on the locked cycle test results, for projecting plant performance.

Figure 18-2: Mesogene Cu Recovery vs Head Grade

Overall copper recovery from the Supergene ore was variable in the samples tested due to the variable proportions of native copper. Native copper is not readily recovered by flotation but can be recovered by gravity concentration as shown by Knelson’s limited test work. There was, however, a reasonable correlation between the floatable copper and total copper grade in the ore as shown in Figure 18-3.

Figure 18-3: Supergene Flotation Copper Recovery vs Head Grade

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Gold Recovery

The Amtel mineralogical study concluded that gold is primarily associated with copper sulphides that will be recoverable by flotation in all 3 ore types. For Supergene, only about 4% of the gold was associated with native copper. For Hypogene ore type, the gold recovery did not correlate well with copper recovery but the gold concentration ratio correlated well with copper concentration ratio as shown in Figure 18-4. Its gold recovery model has been calculated using the concentration ratio correlation and the calculated concentrate production based on copper.

Figure 18-4 Hypogene Au and Cu Concentration Ratio

For Mesogene and Supergene, there was a reasonable direct correlation between gold recovery and copper recovery as shown in Figure 18-5 and Figure 18-6 respectively.

Figure 18-5: Mesogene Au Recovery vs Cu Recovery

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Figure 18-6: Supergene Au Recovery vs Cu Recovery

Silver Recovery

For Hypogene and Mesogene, there was a poor direct correlation of silver recovery with copper recovery, but the silver concentration ratio correlated well with copper concentration ratio. Silver recovery from these ores has been projected using the concentration ratio correlation and the calculated concentrate production based on copper.

Arsenic and Mercury Content in Concentrate

For Hypogene and Mesogene, the flotation test data showed strong correlations between the arsenic and mercury concentration ratios and the copper concentration ratio. These correlations have been used to determine the economical concentrate grade, particularly for Mesogene 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.

18.3 Process Description

The NGD mill was designed to process a blend of underground primary, transition and supergene ores. The process will utilize conventional techniques in progressive particle size reductions, physical separation and concentration of minerals from gangue into concentrates at marketable grades. Particle size reductions will be achieved with crushing, and two (2) stages of primary grinding, while mineral separations will be achieved by gravity concentration and differential flotation. A regrinding stage will be included in the flotation circuit. Processing will not involve chemical processes or alteration of the products. The low concentrations of reagents added during processing will largely be absorbed on particle surfaces to facilitate the physical separation and will not generate chemical reactions.

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The mill was designed to process 11,000 tpd of ore at full capacity and will operate 24 hours per day, 365 days per year with two observed holidays and scheduled downtime for maintenance of equipment. During the first two years of development and operation, the SAG mill will operate as a single grinding stage at a daily rate of 4,400 tpd. The full production rate (11,000 tpd) will be achieved with the completion of the underground to surface conveyor facilities and installation of a Ball mill as a secondary grinding stage. The initial operation is referred to as Phase 1 and then the expanded operation is referred to as Phase 2. The plant will consist of the following unit operations.

• Primary crushing (underground Phase 2);

• Conveying (underground to surface);

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

• SAG Mill primary grinding;

• Ball Mill secondary grinding (Phase 2);

• Closed circuit cyclone classification;

• Gravity concentration;

• Cone crushing SAG pebble recycle;

• Rougher flotation;

• Rougher concentrate regrinding, classification and gravity concentration;

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

• Concentrate thickening and storage;

• Concentrate pressure filtration;

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

• Tailings handling; and

• Water reclaim from tailings.

Figure 18-7 and Figure 18-8 show simplified flowsheets for the Phase 2 process plant.

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Figure 18-7: Crushing and Grinding Simplified Flowsheet

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Figure 18-8: Flotation and Dewatering Simplified Flowsheet

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Coarse Ore Storage and Reclaim

The ore will be crushed underground and conveyed to the surface. Prior to discharge onto a 10,000 tonne live load coarse ore stockpile, there will be a transfer tower with provision to add a waste diversion conveyor to a separate stockpile. (The waste diversion conveyor is not included in the project scope and capital cost estimate). Beneath the coarse ore stockpile there will be a single reclaim tunnel with two apron feeders and a conveyor belt. A dust collector system will be installed in the reclaim tunnel to collect dust generated at the apron feeder loading points and reintroduce the collected fines back onto the SAG Mill conveyor belt.

Grinding

During the start up years, a single stage, SAG Mill will reduce the 150 mm coarse ore to a product size of 80 percent passing 160 microns. A jaw crusher will be temporarily employed on the surface to prepare the mill feed. At full production, a two stage, SAG Mill and a Ball Mill circuit will produce the same product size. During the initial years, the SAG Mill product will be classified in two stages. The SAG discharge screen oversize will be reduced in a Cone Crusher and then recycled to the SAG feed conveyor. The screen undersize will be classified in a bank of cyclones. The cyclone overflow will feed the rougher flotation cells. The cyclone underflow will be returned to the SAG mill feed chute. A slurry stream will be tapped from the cyclone feed to feed the XD-40 Knelson gravity concentrator that will be installed to optimize gold recovery. Knelson concentrates will be directed to the concentrate thickener. Process water additions to the SAG and ball mill circuits will be controlled to maintain densities in the SAG mill, ball mill and the cyclone feed streams.

Flotation

At full tonnage, there will be two parallel banks of flotation cells for the roughing and cleaning applications. One half of the number of flotation cells will be installed initially for the Phase 1 process plant. The primary cyclone overflow will be directed to the flotation circuit. All flotation separations will be conducted at the natural pH of the ore. The collector, and a frother will be added at the beginning of the rougher flotation process. Rougher concentrates will be reground and upgraded to a final concentrate grade with three stages of cleaner and one stage of cleaner-scavenger flotation. A second collector reagent, will be added to the regrind mill to optimize the metallurgical performance in the cleaner circuit.

Rougher concentrate will be reground in a vertical grinding mill operating in closed circuit with cyclones. The cyclone overflow at 80% passing 44 ¼m will be advanced to the cleaning stages.

The first cleaner will operate 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.

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Dewatering and Concentrate Transportation

Concentrate dewatering circuit will involve thickening and pressure filtration, and will consist of the following:

• Concentrate thickener;

• Concentrate stock tank; and

• Pressure filter.

The final concentrates from the third cleaner flotation stage will be dewatered in a thickener. A coagulant and a flocculant will be added to thickener feed to enhance solids settling and to achieve the target underflow density of 55% solids by weight. The thickener overflow will be re-used as process water in the plant via a process water tank.

The underflow from the concentrate thickener will be pumped to a stock tank designed for 24 hour storage. Stock tank concentrate will be pumped to a pressure filter with 40 square meters of filter area, that will produce a concentrate cake containing 8 to 10 % moisture by weight. Concentrates will be sampled, loaded into B-Line trucks and weighed prior to shipment from the site.

Tailings Disposal

The tailings handling circuit will consist of the following unit operations:

• Primary Cyclones (located in the mill);

• Secondary Cyclones (skid mounted @ 45⁰);

• Surface Runoff Pond;

• Pothook Pit;

• Tailings Pond; and

• Sands Drainage Pond.

During the first phase of the mine operation (4,400 tpd), rougher and first cleaner-scavenger tailings will flow by gravity to the final tailings pumpbox. Surface runoff pond slurry will be pumped to the final tailings pump box. The combined tailings will be pumped to the Pothook Pit for containment. Water displaced by the tailings from Pothook Pit will be reclaimed for use in the process plant.

When the plant is expanded (11,000 tpd) the tailings handling system will be modified to deliver two products to the new tailings area. Tailings from the rougher flotation circuit will be pumped to a set of primary cyclones. The 508 mm diameter primary tailings cyclones, located on the flotation cell operating floor, will separate the combined final tailings slurry into slime and sand fractions. Both products will flow by gravity to their respective pump boxes, located outside of, and immediately beside the mill building. The pumps for both sumps will be located inside of the mill building. The cyclone underflow product, at 68% solids, will be diluted with mine and process water, to approximately 30% solids.

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Tailings from the cleaner flotation circuit will be pumped to the primary cyclone overflow pumpbox. The cleaner flotation tailings will contain the sulphide minerals not recovered to the final concentrate. The blend of the cleaner tails and cyclone overflow will be pumped to the pond in the tailings storage facility to prevent sulphides from being used for the construction of the dam walls.

Both cyclone products will be pumped in separate lines in a southerly direction around the west and south of the former Afton mine pit. The cyclone underflow line will arrive at a concrete pier mounted diverter, equipped with two Clarkson type valves. The cyclone underflow lines will then travel in an easterly or in a southerly direction and each branch line will terminate at a skid mounted, 45 degree inclined 508 mm diameter, secondary cyclone pack. The secondary cyclones will be located on the top of the tailings dam. The sand 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.

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19. Mineral Resource and Reserve Estimation (David Rennie, P. Eng., of Scott

Wilson RPA is responsible for sections 19.1 to 19.9 inclusive. Mike Thomas MAusIMM (CP), of AMC Consultants is responsible for section 19.10. )

19.1 Introduction

Scott Wilson RPA has estimated the Mineral Resources for the Project, using a block model constrained by wireframe models of the principal mineralized zones. Grades for Ag, Au and Cu were interpolated into the blocks using Ordinary Kriging (OK). The wireframe models were constructed by NGD personnel and inspected by Scott Wilson RPA prior to their use in the block model. High grades for all three elements were capped before compositing.

19.2 Wireframe Models

Three-dimensional wireframe models were constructed by NGD personnel for each of the supergene, mesogene, and hypogene zones. The principal criteria used for construction of these models were mineralogical characteristics, in contrast to previous models, which were based on a grade shell approach. The boundary of the hypogene zone was defined as the point where chalcopyrite or bornite appeared in the drilling, and pyrite and magnetite (which form a halo around the deposit) disappeared. The mesogene zone is characterized by the appearance of chalcocite, replacing the chalcopyrite and bornite. The supergene zone boundary is defined by the onset of native Cu in the drill core.

Scott Wilson RPA inspected the wireframe models and considers them to represent reasonable and plausible interpretations of the mineralization at New Afton. They encompass a series of reasonably coherent tabular zones that strike EW and dip vertically to moderately northwards. The bulk of the mineralization occupies a single large vertically-dipping slab with portions of each of the three mineralization types (see Figure 10-1).

These models were assigned integer codes as follows:

• Hypogene 2001

• Mesogene 2002

• Supergene 2003

The wireframes were used to assign these codes to drill hole composites and blocks within the block model. This allowed the computer to discriminate between blocks and composites within each of the zones for the purposes of grade interpolation and rock type assignment (which, in turn, dictated the bulk density to use). The wireframes were also used to estimate the percentage of each block contained within each wireframe. Separate percentage models were generated for each mineralized zone, and then used to create a combined block model. The percentage models provide more accurate tonnage estimates because they allow for accumulation of volume from partial blocks.

Scott Wilson RPA notes that where the wireframes meet, there are small gaps and overlaps. This created some errors in the percentage model calculation, as when the percentages were totalled, some blocks reported more or less than 100%. In these instances, the total percentage was adjusted to 100% to ensure that volumetrics calculations were not adversely impacted.

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19.3 Sample Database

The samples used for the estimate were derived from diamond drilling. A total of 8,566 samples, out of a database comprising 22,926 sampled intervals, were contained within the wireframe models. Not all the samples have been assayed for all economic elements. Non-sampled or missing intervals were assigned zero grade prior to compositing.

The database contained information for holes up to and including UA-66. The assay information was compiled by NGD into spreadsheets which were then imported to GEMS. Scott Wilson RPA validated the database by comparing a portion of the assays against the original lab reports. A total of 12,222 samples, representing 53% of the sample database were compared to the lab reports. There were very few discrepancies noted, and no systematic errors were found that would impact estimation of Mineral Resources In addition to this check, Scott Wilson RPA applied the GEMS validation utilities to the database in order to check that the sample intervals were entered correctly. A modest number of discrepancies were found which appeared to be key-punching errors, primarily in the recording of from and to intervals. These errors or inconsistencies were flagged by Scott Wilson RPA and corrected by NGD personnel.

Scott Wilson RPA checked the lithology, collar coordinates and downhole surveys recorded in the database against the original paper logs for approximately 12% of the drill holes (selected more or less at random). These checks focused primarily on the underground drilling that commenced in 2005, but also included a few of the pre-2005 holes. There were no errors or discrepancies noted.

In Scott Wilson RPA’s opinion, the drilling database for the New Afton project is relatively free from errors and is acceptable for use in estimation of Mineral Resources.

19.4 Capping of High Grades

Sample statistics are presented in Table 19-1.

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Table 19-1: Sample Statistics

Hypogene	Ag	Au	Cu
Num	3936	4136	4257
Mean	1.977	0.892	1.102
Median	1.700	0.690	0.960
SD	1.828	0.769	0.870
CV	0.925	0.862	0.790
Mesogene	Ag	Au	Cu
Num	3315	3365	3584
Mean	4.196	0.929	1.203
Median	2.200	0.490	0.860
SD	5.348	1.312	1.162
CV	1.274	1.412	0.966
Supergene	Ag	Au	Cu
Num	610	606	653
Mean	4.053	0.749	1.224
Median	2.300	0.340	0.840
SD	4.656	1.401	1.177
CV	1.149	1.871	0.962

The grade distributions for all components are positively skewed which, in Scott Wilson RPA’s opinion, creates a potential for overestimation of the average grade of the deposit. This requires that some technique should be applied to moderate the affect of the high-grade “tails” of the distributions. Scott Wilson RPA capped high grades to the values listed in Table 19-2.

Table 19-2: Cutting Values

		Cap			Num. Cut
	Ag	Au	Cu	Ag	Au	Cu
	(g/t)	(g/t)	(%)
Hypogene	10.0	3.4	3.5	10	42	46
Mesogene	25.0	4.5	4.0	29	85	102
Supergene	22.0	4.5	4.0	5	13	21

Selection of capping values is difficult in the absence of production experience. The cap values were chosen based on inspection of the histograms and probability plots, and on “cutting curves”. Cutting curves are graphs of the mean sample values at a range of caps. The capped mean is observed to diminish with increasingly lower cap values, and the cutting curve becomes progressively steeper as it approaches the Y-axis. In Scott Wilson RPA’s opinion, a reasonable cap value is generally found in the flatter portion of the curve just to the right of where it steepens towards the origin.

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19.5 Compositing

Samples were composited to 4 m prior to grade interpolation. Ideally, the composite length should be smaller than the block size but at least as large as the longest sample. Scott Wilson RPA reviewed the lengths for the samples contained within the wireframe models and notes that they range from 0.2 m to 9.14 m. The most commonly used sample length was 2m, but it is reported that this standard for the sampling was established in 2005. Scott Wilson RPA notes that approximately 97.5% of the samples are less than or equal to 4 m long. In Scott Wilson RPA’s opinion, compositing to the longest sample length (i. e., 9.14 m) would be impractical due to the block size (10 m x 10 m x 10 m). As noted above the majority of samples are less than or equal to 4 m in length so that was chosen as the compositing interval.

Samples were composited to 4 m downhole lengths, with no breaks for lithological boundaries. Scott Wilson RPA notes that this will result in the addition of as much as 2 m of dilution at the boundaries of the wireframe models. However, in Scott Wilson RPA’s opinion, the amount of dilution added will be small relative to the tonnage of the Mineral Resource. Also, the boundary of the deposit is not particularly sharp, so it is not unreasonable to include a modest amount of dilution on the margins.

Statistics for the 5,027 composites contained within the wireframe models are shown in Table 19-3.

Scott Wilson RPA notes that there is a significant difference between the declustered and non-declustered means. This suggests that there is clustering of composites in higher-grade areas of the deposit, which can result in overstatement of the Mineral Resource grade if not properly handled. For this estimate, the problem of clustering was dealt with by virtue of the fact that kriging was used for grade interpolation. Kriging declusters the data, so in Scott Wilson RPA’s opinion, there should not be a problem with bias due to clustering.

Scott Wilson RPA further notes that there are trends evident in the mean grades for the composites. For example, the Ag grade is observed to be much higher in the mesogene and supergene zones compared to the hypogene zone, which suggests that some secondary enrichment has taken place. The same trend is apparent for Cu but it is not nearly as pronounced as for Ag. The opposite trend occurs for Au, suggesting that there may be depletion of depletion of Au in the weathered zones in the weathered zones.

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Table 19-3: Composite Statistics

Hypogene	Ag	Au	Cu
Num	2361	2361	2361
Mean	1.694	0.810	1.028
Decl. Mean	1.437	0.662	0.852
Median	1.100	0.431	0.680
SD	1.356	0.649	0.752
CV	0.943	0.980	0.883
Mesogene	Ag	Au	Cu
Num	2288	2288	2288
Mean	3.822	0.820	1.158
Decl. Mean	3.038	0.672	0.863
Median	1.442	0.376	0.530
SD	3.953	0.794	0.902
CV	1.302	1.180	1.045
Supergene	Ag	Au	Cu
Num	378	378	378
Mean	3.782	0.609	1.155
Decl. Mean	3.108	0.502	0.964
Median	1.733	0.245	0.611
SD	3.520	0.669	0.885
CV	1.133	1.332	0.919

19.6 Bulk Density

NGD performed a total of 436 density determinations on drill core specimens. These determinations were carried out at the Eco-Tech lab using the water immersion method on wax-sealed core samples. Scott Wilson RPA conducted statistical analyses on the density data, grouped by mineralized zone. The mean value of the density samples for each zone was applied to the blocks within these zones to estimate the tonnages. The values used were, 2.61 t/m³ for the hypogene and mesogene zones, and 2.57 t/m³ for the supergene zone.

19.7 Geostatistics

Scott Wilson RPA carried out geostatistical analyses for Ag, Au, and Cu, in each of the hypogene, mesogene, and supergene zones. The analyses comprised 3-D and linear pair-wise relative semi-variograms created using GSLIB software. Pair-wise relative semi-variograms normalize the difference between each sample pair by the mean of that pair. These variograms are resistant to variability in the data and often show trends where conventional semi-variograms do not. The semi-variograms were then used to derive variogram models for kriging.

The longest ranges for semi-variogram models for all components tended to occupy a planar zone parallel to the principal strike and dip of the deposit (i.e., striking EW with a near vertical dip). The variography for the hypogene zone was typified by a strongly defined plunge at approximately 45⁰ downwards toward grid west. This trend was not apparent for the mesogene and supergene zones, further suggesting that weathering processes have altered the distribution of the principal economic elements.

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In general, the semi-variograms were more coherent for the hypogene zone than for the mesogene or supergene zones. In all zones, and for most elements, the major axes have ranges measuring 100 m or more. However, Scott Wilson RPA notes that for most of the variogram models in the meso- and supergene zones, 2/3 to ¾ of the sill value was achieved within a range of 20 – 30 m. This is particularly true for Au in the supergene zone and Cu in the mesogene.

19.8 Block Models

The block models were constructed in GEMS Ver. 6.0. Block size was 10 m x 10 m x 10 m, and there was no rotation applied (i. e., the edges of the block model parallel the mine grid¹ NS and EW directions). Percentage models were used to account for partial blocks, so no sub-blocking was required.

Separate block models were constructed for each of the mineralized zones. Grades from these three models were combined into a single master block model. The grades in the master block model were calculated by using the volume-weighted average of grades from each of the other models. Block rock code assignments were made using the wireframe models. A code was assigned to a block if any part of it was contained within a particular wireframe. Blocks contained within more than one wireframe were assigned a code based on a hierarchy which assigned highest priority to hypogene, followed by mesogene, and finally supergene. This meant that rock codes in the master block model might not match those in the individual zone models. However, this has no real impact on the grade estimates.

Block model geometry is listed in Table 19-4 below. The GEMS convention for the block model origin is that it is at the lower left (SE) corner, at the top of the block model. Levels are counted downwards from the top level. The origin is at the corner of the first block, not the centroid.

Table 19-4: Block Model Geometry

Origin:
	X	2800
	Y	1400
	Z	5750
Size:	10 m x 10 m x 10 m
Columns:	160
Rows:	120
Level:	115
Rotation:	None

19.8.1 Search and Kriging Parameters

Search ellipsoids were oriented the same as the variogram models for each element and domain (zone). The search ranges were also made the same as the variogram models out to a maximum of 200 m. Each estimate was limited to a minimum of 1 and a maximum of 12 composites, with a maximum of 3 allowed from any one drill hole. Block grades were calculated from an average of 27 point estimates (i. e., 3x3x3 block discretization).

¹ The orientation of the mine grid relative to the UTM system is explained in section 25.1

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19.8.2 Cut-Off Grades

Following metallurgical testwork, the recoveries listed in Table 19-5 were used to prepare a resource block model for use in estimating the mineral reserve.

Table 19-5: Metal Price and Recovery Factors

Recoveries
	Cu	Au	Ag
Hypogene	93	87	75
Mesogene	89	81	75
Supergene	85	80	75
Metal Prices
Cu		US$1.20/lb
Au		US$450/oz
Ag		US$5.25/oz
Exchange		US$0.88/C$1.00

The Dollar Value equation was as follows: 

DolVal = (Cu x 1.2 x Cu_Rec x 22.046) 

+ (Au x 450.00 x Au_Rec/ 31.103) 

+ (Ag x 5.25 x Ag_Rec / 31.103))

Recovery values were varied depending on the block rock code.

The Dollar Value/t was used for application of a cut-off grade. A cut-off grade of C$10/t was applied to the block model for the Mineral Resource estimate.

19.8.3 Block Model Results

The unclassified block model results at a range of Dollar Value cut-offs is provided below in Table 19-6.

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Table 19-6: Block Model Results

Cut-off	Volume	Density	Tonnage	Cu	Au	Ag	DOLVAL
C$/t	m3	t per m3	Kt	(%)	(g/t)	(g/t)	C$/t
$15.00	22,500	2.606	58,644	1.102	0.827	2.730	36.79
$10.00	25,195	2.606	65,659	1.024	0.771	2.591	34.21
Total	26,555	2.606	69,201	0.983	0.741	2.512	32.85

A tonnage curve for various cut-offs is shown in Figure 19-1.

Figure 19-1: Tonnage Curve

19.8.4 Block Model Validation

The block model was subjected to the following validation exercises to check the grade interpolation:

• Inspection of block grades in plan and section and comparison with drill hole grades.

• Cross-validation (“jack-knifing”).

• Comparison of global mean block grades and composite grades.

• Comparison with a grade interpolation done using an Inverse-Distance-Squared (ID2) weighting.

Block grades for all three elements were examined exhaustively for each column, row and level in the model and compared to the drill hole composites. The block grades are observed to honour the composite grades quite well.

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Cross validation, or jack-knifing was used to check the kriging model and search parameters. The cross validation process involves sequentially removing each data point and estimating the value of that sample using the surrounding samples. A comparison of the mean composite grades with those estimated during the cross validation process is provided in Table 19-7.

Table 19-7: Cross-Validation Results

Hypogene	Data	Estimate	Diff	% Diff
Ag	1.702	1.694	-0.01	-0.5	%
Au	0.822	0.810	-0.01	-1.5	%
Cu	1.044	1.028	-0.02	-1.5	%
Mesogene	Data	Estimate	Diff	% Diff
Ag	3.775	3.822	0.05	1.2	%
Au	0.845	0.820	-0.03	-3.0	%
Cu	1.166	1.158	-0.01	-0.7	%
Supergene	Data	Estimate	Diff	% Diff
Ag	3.881	3.782	-0.10	-2.6	%
Au	0.613	0.609	0.00	-0.7	%
Cu	1.167	1.155	-0.01	-1.0	%

The global means of point estimates compare reasonably well with the data means. Global block means are compared with the composite means in Table 19-8.

Table 19-8: Comparison of Block and Declustered Composite Grades

Hypogene	Comps	Blocks	Diff	% Diff
Ag	1.437	1.457	0.02	1.4	%
Au	0.662	0.709	0.05	7.1	%
Cu	0.852	0.882	0.03	3.5	%
Mesogene	Comps	Blocks	Diff	% Diff
Ag	3.038	3.337	0.30	9.8	%
Au	0.672	0.727	0.05	8.2	%
Cu	0.863	0.950	0.09	10.1	%
Supergene	Comps	Blocks	Diff	% Diff
Ag	3.108	3.122	0.01	0.5	%
Au	0.502	0.483	-0.02	-3.8	%
Cu	0.964	0.985	0.02	2.2	%

The global block means agree reasonably well with the declustered composite means for all components.

The mean block Cu grade is noted to be 10% higher than the declustered composite mean, which in Scott Wilson RPA’s opinion, is just within an acceptable margin. Scott Wilson RPA further notes that there appears to be a positive bias between the block grades and data grades for the hypogene and mesogene zones (i.e., block grades are higher than the composite grades).

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A grade interpolation was run using ID2 weighting. This interpolation used the same search parameters as the kriged estimate. A comparison of the ID2 and kriged estimates is provided in Table 19-9. The comparison was done using a Cu equivalent (Cueq) grade determined by taking the Dollar Value and back-calculating a grade using the Cu recovery and metal price. The cut-off for generation of tonnage curves for these estimates was based on the CuEq grade.

Table 19-9: Comparison of Kriged and ID2 Estimates

Interpolation	Tonnage		Cu		Au		Ag		CUEQ
Type	Kt		(%)		(g/t)		(g/t)		(%)
Kriging	69,198		0.983		0.741		2.512		1.371
ID2	69,198		1.011		0.746		2.584		1.402
Difference	0.000		0.029		0.004		0.072		0.031
% Diff.	0.0	%	2.9	%	0.5	%	2.9	%	2.3	%

Note: Tonnage totals differ from Table 19-15, because this comparison was run on a preliminary model.

As can be seen in Table 19-9, the ID2 estimate compared quite closely to the kriged estimate on a global basis. The graph in Figure 19-2 shows the tonnage curves for each estimate at a range of CuEq cut-offs. The ID2 estimate is observed to have a higher proportion of material at higher cut-offs and a lower proportion of lower grade material than the kriged estimate. In Scott Wilson RPA’s opinion, this implies that kriging has smoothed the grade distribution somewhat more than the ID2.

Figure 19-2: Tonnage Curves For ID2 VS OK

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Grade smoothing is, in part, a function of the block size of the model. Smoothing lowers the variance of the grade distribution, reducing the number of extreme values and in effect, drawing in the ends of the grade histogram. The larger the block size, the more smoothing (i.e., variance reduction) that occurs. When a cut-off is applied to any set of blocks, if the cut-off is above the mean grade of the deposit, then as the block size goes up (i. e., the variance goes down), the amount of material reported above the cut-off is reduced. The opposite is true for cut-offs below the mean. In this case, the reduction in variance results in more material reporting above the cut-off than is predicted by the block model. If possible, the block size of the model should be calibrated to the expected size of the selective mining unit (SMU) to ensure that it will report the Mineral Resources that will actually be mined.

The mining method proposed for the deposit is block caving. The block caving method does not afford much grade selectivity once the caving has been initiated. In Scott Wilson RPA’s opinion, grade smoothing in the model is probably somewhat low relative to the block caving selectivity. The cut-off proposed for the project is below the mean for the entire block model (i.e., $10/t vs a mean value of $32.85/t), which suggests that there is more material above cut-off in the deposit than the model reports. This suggests that the block model is conservative in reporting Mineral Resources, and that there should be no negative impacts from the degree of smoothing in the block model.

19.8.5 Dilution Halo

Preliminary mine design and simulations based on a block caving mining method resulted in the definition of a “cave volume” encompassing the planned mined material. In addition to this wireframe model, a dilution halo was added to the planned cave volume. The dilution halo extends for a distance of approximately 30 m on the sides and top of the planned cave. The material contained within both the planned cave volume and the dilution halo consists of resource blocks (i.e., is contained with the resource wireframe models) as well as waste material outside of the Mineral Resource wireframe boundary. Some of this waste material contains significant mineralization. The mineralized portions of the dilution halo were not included in the Mineral Resource volume models because they were either too remote from the main body of mineralization, or they could not be included due to local structural complications in the interpretation. However, in Scott Wilson RPA’s opinion, it is reasonable to assume that the material immediately surrounding the Mineral Resources is not zero-grade.

To assist in simulation of the mine production, Scott Wilson RPA was instructed to prepare a grade estimate of the diluting material outside of the Mineral Resource volume. This model was configured as an add-on to the existing block model created for the Mineral Resource estimate. The block grade estimates are only preliminary in nature owing primarily to the scarcity of data in the dilution halo . No dilution has been included in the Mineral Resource estimate.

The dilution blocks and composites were assigned an integer code. Composites were coded based on centroid location relative to the resource model solids. Generally, anything outside of the mineral resource wireframes were allowed to be used in the dilution estimate. Composites already coded as within the resource volume (i.e., codes 2001, 2002, or 2003) were left unchanged. The grade interpolation was carried out using only composites coded as “Dilution”.

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The dilution halo composites were extracted and subject to statistical and variogram analyses for Ag, Au, and Cu. The composites were observed to contain relatively high numbers of zero values, as much as 30% of the database for Au and Ag. The grade distributions of the remaining composites were observed to be moderately to highly skewed, similar to those of the resource composites. In Scott Wilson RPA’s opinion, there is a risk of overestimation of the average grade of the blocks unless some technique is applied to moderate the affect of the high-grade “tails” of the distributions. Capping of the sample grades was done for the resource model; however, the cap values used for resource composites were not appropriate for the dilution halo. It was not possible to apply a different set of cap levels to the dilution samples without reconstructing the existing composite database. For expedience, it was decided to restrict the range of influence of high-grade composites.

Probability plots of the composites showed that there were high-grade populations for each of Ag, Au, and Cu. The lower thresholds for these populations, estimated from the probability plots, were roughly 2 g/t Ag, 0.5 g/t Au, and 0.6% Cu. Omni-directional indicator variograms were run for each of these threshold values, and all yielded a range of 29 to 30 m. A limit of 30 m was placed on the range of influence of all composites above the threshold values.

Geostatistical analyses yielded a different set of anisotropy axis orientations from the resource model. All three semi-variogram models were oriented as vertical planes striking grid north, with major axes all plunging vertically. Major axes had ranges of 130 m for Ag, 155 m for Au, and 120 m for Cu. Semi-major axes (all oriented grid NS) were 55 m for Ag, 60 m for Au, and 55 m for Cu. Minor axes measured in the order of 20 m for all components. This orientation is roughly perpendicular to the resource body.

The variogram model orientations do not match any known geological feature of the deposit, which in Scott Wilson RPA’s opinion, is somewhat unexpected. The fracture systems that are thought to control the resource mineralization also occur in the walls of the deposit and should therefore have had some influence on the dilution grades. Scott Wilson RPA recommends that additional geological interpretation be carried out to try and resolve the controls on mineralization outside of the resource body.

The geostatistics did not support the present geological interpretation, so the search parameters were not configured by the variography. Instead, a spherical 100 m search was used, based loosely on the average range of the variograms and deliberately configured to work well with the 80 m drill section spacing outside of the resource body. Block estimates were carried out by Ordinary Kriging, with a minimum of one, maximum of 12 comps, and a maximum of 3 comps from any one hole. These are the same as the resource model parameters. No classification was applied to the dilution blocks, as they were not intended to be included in the Mineral Resources.

The resulting block model was combined with the resource model. Grades for blocks containing both resource and dilution were estimated using an average weighted by the relative proportion of resource and/or dilution material in each block. This model was then exported for further mine simulation and design.

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Scott Wilson RPA notes that there are several deficiencies in the dilution halo model. The drill holes traversing the diluting material were not sampled as rigorously as the mineralized body. Consequently, there are several gaps in the sampling, and numerous intervals for which it was necessary to assign zero grade owing to lack of sample data. The grade interpolation was unconstrained, save for the limit placed on the range of influence of high-grade composites. This has resulted in smearing of grades across fairly large volumes, in a pattern that is more dictated by the search ellipsoid than any geological constraint. The variography yielded models that are orientated at almost right angles to any known geological structures including the strike and dip the resource body. In Scott Wilson RPA’s opinion, the dilution halo model provides only a preliminary estimate of the grades outside of the resource volume and cannot be included in the Mineral Resource estimate.

19.8.6 Arsenic and Mercury Model

Preliminary metallurgical results indicated that there were elevated levels of Hg and As in some areas of the deposit, particularly in the supergene and mesogene zones. Scott Wilson RPA was instructed to prepare a preliminary block model estimate of the As and Hg content within the proposed cave volume. The purpose for the block model was to provide a means to try to predict the As-Hg grades in ore produced during the LOM.

The database for As and Hg is very small relative to the database for the economic elements. NGD personnel selected a suite of 1,125 pulps from the drill samples to submit for ICP assay for a suite of elements which included Hg and As. The samples were collected from every 5^th sample in the drill holes (i. e., every 5 m), from a variety of localities within and surrounding the deposit but principally in the supergene and mesogene zones. Scott Wilson RPA notes that the density of data for As and Hg is very much lower than for the economic elements which significantly reduced the confidence level of the block model grade estimates.

Scott Wilson RPA carried out statistical analyses on the As-Hg data. The analyses included general statistics, histograms, probability plots, and correlation coefficients. A statistical summary is provided below in Table 19-10.

As and Hg values are clearly elevated for many of the samples. The statistics show highest average mean and median As and Hg grades in the mesogene and supergene zones. The grade distributions are highly skewed, in a similar fashion to the economic metals. A review of the correlation coefficients indicates that As and Hg show very strong positive correlations with each other as well as with Sb (not included in the model). However, there is only a modest correlation between As and both Cu and Ag, and none between the economic elements and Hg. No significant relationship was evident between Au and either Hg or As. The probability plots demonstrated that there are multiple sample populations in each of the ore-type sub-groups, which implies that sub-setting the data by ore zone is not particularly meaningful.

In Scott Wilson RPA’s opinion, the statistics suggest that the As and Hg have little or no genetic association with the economic mineralization, even though they are spatially related. That both deposits happen to occur in the same place is probably more a result of the structural controls of both mineralizing events than any genetic relationship. The faults and fractures through the deposit area probably acted as conduits for fluids for both processes, resulting in a complex overprinting of different mineral species within a fairly restricted structural corridor.

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Table 19-10: As-Hg Sample Statistics

As (ppm)	All	Hypogene	Mesogene	Supergene	Waste
Num.	1125	31	676	94	324
Mean	372.1	132.9	460.2	609.7	142.3
SD	1073.6	373.1	1251.1	1242.3	429.6
CV	2.89	2.81	2.72	2.04	3.02
Median	40.0	20.0	45.0	82.5	30.0
Max	10000.0	1925.0	10000.0	7010.0	5110.0
Min	2.5	2.5	2.5	2.5	2.5
Hg (ppb)	All	Hypogene	Mesogene	Supergene	Waste
Num.	1125	31	676	94	324
Mean	2916.3	425.9	3054.3	7374.0	1573.4
SD	10260.1	980.0	10060.9	19665.7	5816.3
CV	3.52	2.30	3.29	2.67	3.70
Median	209.0	100.0	246.0	176.0	125.0
Max	146000.0	4960.0	124000.0	146000.0	72000.0
Min	2.5	26.0	2.5	20.0	2.5

NGD personnel constructed a wireframe model around what appeared to be the higher-grade component of the As-Hg mineralization. This body consisted of a lens, partially encompassed by the supergene and mesogene solids, bounded by the HW Fault and a sub-parallel splay called the Picrite Fault. Statistics for the database segregated by this wireframe model (termed “PC zone”) are shown in Table 19-11.

Table 19-11: As-Hg Sample Statistics – Wireframe

As (ppm)	All	PC Zone	Outside
Num.	1125	760	365
Mean	372.1	514.6	75.3
SD	1073.6	1272.9	219.9
CV	2.89	2.47	2.92
Median	40.0	70.0	20.0
Max	10000.0	10000.0	1925.0
Min	2.5	2.5	2.5
Hg (ppb)	All	Wireframe	Outside
Num.	1125	760	365
Mean	2916.3	4126.4	395.6
SD	10260.1	12273.7	1180.3
CV	3.52	2.97	2.98
Median	209.0	475.0	84.0
Max	146000.0	146000.0	12800.0
Min	2.5	8.0	2.5

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The samples within the wireframe are clearly higher in grade on average. Splitting the database into two groups using the PC wireframe yielded somewhat simpler grade distributions for both As and Hg, suggesting that this model is better suited to constraining the grade interpolation than the “xgene” solids (see Table 19-12). The PC wireframe provided a rough constraint for the grade interpolation but it could probably be improved. Scott Wilson RPA recommends that routine analyses for As and Hg be incorporated in the sampling protocols at New Afton in order to improve both the geological understanding and the quality of the grade interpolations. It is also recommended that once enough data has been collected, that geological interpretation work be carried out to resolve the controls to As and Hg mineralization so that appropriate modeling constraints can be developed.

All of the data are highly skewed, which suggests some means of moderating the effect of the higher grade tails of the distributions should be applied during grade interpolation. The range of influence of higher-grade samples were limited to 10 m (i.e., one block) for all samples with grades greater than or equal to 6,000 ppm As or 50,000 ppb Hg. These high-grade thresholds were selected from the probability plots. There are distinct high-grade populations within the PC zone samples for both As and Hg. The 6,000 ppm As and 50,000 ppb Hg values are the lowermost thresholds for these high-grade populations.

The data could not be composited owing to the sampling strategy of selecting a 1 m interval every 5 m. In Scott Wilson RPA’s opinion, this will result in higher variability in the sampling, and a higher degree of grade smoothing in the model.

Geostatistical analyses were carried out for As and Hg. The axes directions of the variogram models were somewhat similar to those of Cu, Au, and Ag. Both As and Hg had major axes oriented at grid azimuth 000⁰ and plunging upwards at 60⁰, with the semi-major axis pointing horizontally grid east. The models represent ellipsoids with a grid-EW strike, dipping moderately to the south. This is similar in strike but slightly different in dip to the models for the economic elements in the resource volume (see section of this report entitled Geostatistics). Total ranges were 140 m x 60 m x 50 m for As and 160 m x 60 m x 60 m for Hg, although it should be noted that both models required two structures and that the 90% of the total sill for As was attained at ranges of 40 m x 59 m x 20 m (the range of the 1^st structure).

The grade interpolation was carried out using OK and a search ellipsoid measuring 65 m x 60 m x 40 m, oriented parallel to the principal axes of the variogram models. Block estimates were limited to a minimum of 2 and maximum of 12 samples, with a maximum of 3 allowed from any one drill hole. Grade estimates were constrained to blocks wholly or partially enclosed by the dilution halo wireframe model (see section of this report entitled Dilution Halo).

There were many blocks within the dilution halo that did not receive a grade estimate due to scarcity of data points. For these blocks the global median grades of the data sets were applied. These values were 70 ppm As and 475 ppb Hg for inside the PC wireframe and 20 ppm As and 84 ppb Hg outside. The median value was used instead of the mean because the data were so highly skewed that, in Scott Wilson RPA’s opinion, the mean would overstate the actual grades.

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In Scott Wilson RPA’s opinion, the confidence level of the grade interpolation for As and Hg is very low. This is due primarily to the relative lack of assay data. This has made interpretation and geostistical analysis difficult, which has in turn, made it difficult to properly configure the grade interpolation. In addition, it was not possible to fill all of the blocks using a reasonable search distance. The unfilled blocks were assigned a median grade for their respective data populations, which in Scott Wilson RPA’s opinion, is a very crude means of estimating grade. Scott Wilson RPA strongly recommends that the As and Hg database be expanded to encompass the entire deposit area so that a more rigorous grade interpolation exercise can be carried out.

19.8.7 Classification

Classification of the Mineral Resource estimate was carried out using the definitions in the CIM Standards on Mineral Resources and Reserves Definitions and Guidelines, to be consistent with the requirements of NI43-101. Mineral Resources are assigned to one of three categories depending on the confidence level of the estimate.

In Scott Wilson RPA’s opinion, if possible, it is desirable to derive a single set of classification parameters that can be applied to the entire deposit. However, this is complicated by the fact that the estimate contains multiple elements for which the grade interpolation parameters and number of data points vary. In addition, the grade estimates were carried out using variogram models unique to each mineralized zone. Individual classification rules could be worked out for each element in each zone, but different classifications could be assigned to the same block depending on the zone and element.

Scott Wilson RPA notes that project economics are most dependent upon the Cu and Au grades. Of these two elements, the variogram ranges and by extension, the estimation confidence levels for Au are observed to be somewhat less than for Cu. Consequently, the principal components used for defining the classification of the deposit were those of the Au estimate. In Scott Wilson RPA’s opinion, if the classification is assigned based on the confidence level of the Au, then the confidence level for Cu estimate should be at least as good. This should ensure that the classification is valid for both the Au and Cu estimates.

The present drill section spacing is a nominal 40 m, which is approximately one half of what it was prior to the start of the 2005 in-fill drilling program. With completion of the in-fill drilling, the general form and location of the deposit did not change much. This demonstrates, in Scott Wilson RPA’s opinion, that there is good geological continuity from section to section and the location of the zone can be predicted from the geological model with a fair degree of certainty. For this reason, Scott Wilson RPA considers all mineralization contained within the interpreted geological boundaries (i. e., the wireframes) to be in at least the Indicated category.

For definition of Measured Resources, Scott Wilson RPA used the distance to the nearest composite. An integer code was assigned to the blocks to signify level of confidence. Initially, all blocks inside the wireframes were given a code of 2, which is analogous to a classification in the Indicated category. Then, all blocks within 45 m of a composite were assigned an integer code of 1 signifying the highest level of confidence. Following this any class 1 blocks intersected by a wireframe boundary (i.e., blocks on the edge of the mineralized body) were downgraded to class 2. Class 1 blocks became Measured, and class 2 blocks were assigned as Indicated.

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The 45 m distance was employed because it is roughly 1/3 of the omni-directional semi-variograms ranges for Au in the hypogene and supergene zones (these ranges were 150 m for hypogene, 65 m for mesogene, and 140 m for supergene). The 45 m distance also works well with the in-fill drill spacing carried out in 2005, such that those portions in the core of the deposit confirmed by the in-fill drilling were generally categorized as Measured.

19.9 Statement of Mineral Resources

The Mineral Resource estimate, published in September 2006, is summarized in Table 19-12.

Table 19-12: Mineral Resources (at US$1.20/lb Cu, US$450 Au and US$5.25 Ag)

Measured
Cut-off	Tonnes	Cu	Au	Ag	DOLVAL
(C$/t)	Kt	(%)	(g/t)	(g/t)	C$/t
$15.00	39,900	1.18	0.87	2.79	$39.35
$10.00	43,250	1.12	0.83	2.68	$37.26
Indicated
Cut-off	Tonnes	Cu	Au	Ag	DOLVAL
(C$/t)	Kt	(%)	(g/t)	(g/t)	C$/t
$15.00	18,800	0.93	0.73	2.60	$31.37
$10.00	22,410	0.84	0.66	2.42	$28.34
Measured & Indicated
Cut-off	Tonnes	Cu	Au	Ag	DOLVAL
(C$/t)	Kt	(%)	(g/t)	(g/t)	C$/t
$15.00	58,700	1.10	0.83	2.73	$36.79
$10.00	65,660	1.02	0.77	2.59	$34.21

* Recovered value, assuming metallurgical recoveries of 90% for Cu and Au, and 75% for Ag, and a C$:US$ Exchange Rate of 0.88

The total Measured and Indicated Mineral Resources are 65.66 Mt at the C$10/t cut-off, with a total metal content of 1,480 million pounds of Cu, 1.6 million ounces of Au and 5.5 million ounces of Ag.

In addition to this estimate, a new estimate of Mineral Resources was carried out for the recently discovered C Zone, which is immediately below the existing resource body (Main Zone). It is categorized as Inferred Mineral Resources. The estimate was performed using the same methodology and parameters as those for the Main Zone. Grades were interpolated using Ordinary Kriging, using the results of 11 underground diamond drill holes and one surface hole. The mineralization is all Hypogene, or primary, with chalcopyrite being the dominant Cu-bearing mineral. A wireframe model, constructed by NGD personnel, was used to constrain the block model. A specific gravity of 2.61 t/m³ was used in the estimation (consistent with that used in the Main Zone estimate).

The C Zone Inferred Resource estimate is summarized in Table 19-14.

Table 19-13: C Zone Inferred Resource At US$1.20 Cu, US$450 Au, and US$5.25 Ag

Cut-Off)	Tonnes	Cu	Au	Ag	DOLVAL
(C$/t)	(Kt)	(%)	(g/t)	(g/t)	(C$/t)
$15.00	6,590	1.10	0.97	1.75	$39.41
$10.00	7,940	0.96	0.88	1.55	$34.89

* Recovered value, assuming metallurgical recoveries of 90% for Cu and Au, and 75% for Ag, and a C$:US$ Exchange Rate of 0.88

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The total Inferred Mineral Resources at the C$10/t cut-off, are 7.94 Mt grading 0.96% Cu, 0.88 g/t Au, and 1.55 g/t Ag, with a total metal content of 168 m lbs Cu and 0.224 m oz Au.

Due to the relatively low density of drilling information, the C Zone has been classified as Inferred. Additional infill drilling will be required to increase the confidence level in the resource and upgrade it to the Indicated and/or Measured categories.

The resources from the C Zone were not used in the estimation of Mineral Reserves in this Technical Report.

19.10 Mineral Reserves

The Mineral Reserves were estimated by Mike Thomas MAusIMM(CP), an employee of AMC Consultants Pty. Ltd and a Qualified Person (under National Instrument 43-101).

19.10.1 Sources of Information

The Mineral Reserve has been estimated using the resource block model prepared by Scott Wilson RPA, the contents of which form the basis of the Mineral Resource Estimate summarized in Table 19.12. To enable AMC to account for the effects of dilution and to estimate arsenic and mercury grades, Scott Wilson RPA provided AMC with a Combined Resource Model constructed by surrounding the mineral resource model with a block model of the halo zone, which in turn was surrounded by blocks containing waste (rock with no mineral values and a density of 2.61 t/m³). Values reflecting the estimated grades of mercury and arsenic were also added to the model.

Geotechnical data and geological interpretations of faults and other structures used in the estimation process have been provided by NGD. Mill recoveries, concentrate grades, concentrate transportation costs, concentrate treatment charges and other information relating to the revenue generated by concentrate sales, including metal prices, have been provided by Hatch. Ore processing and other site costs, other than the costs of mining, have also been provided by Hatch.

AMC has reviewed the information provided by Scott Wilson RPA, NGD and Hatch for major inconsistencies and errors material to its work, but has not audited the geological data collection process, the interpretation of the geological data, or the resource modeling process.

19.10.2 Preparation of the Resource Model for Reserve Estimation

The Combined Resource Model, prepared using GEMS software, was provided to AMC as a text file, which AMC converted to Datamine² format, then added a resource value field and fields containing geotechnical information to facilitate the estimation process.

To enable a single cut-off parameter to be used when selecting the economic parts of the resource to be mined, AMC estimated the theoretical value (NSR Value) received for concentrate produced from each tonne of resource, net of all costs and losses incurred in the off-site transportation and processing of concentrate. Metal prices, metallurgical recoveries and other values used in the estimation are shown in Table 19-14 and Table 19-15.

² Datamine geological modelling software, a product of Datamine Corporate Limited

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It is envisaged that the various geotechnical domains within and surrounding the resource will fragment differently during caving and consequently will move at different rates through the cave mass to the drawpoints. To enable the different fragmentation properties to be modelled, blocks lying within identifiable geotechnical domains were assigned geotechnical domain codes (GTDOM). In addition to the hypogene, mesogene and supergene rock types already identified in the resource model, picrite, waste and halo material were identified and assigned rock type codes (RKTYPE).

After addition of the NSR Value field and the geotechnical and rock type codes, the upper surface of the model was trimmed by the surface topography. The final 10m x 10m x 10m block model contained the fields described in Table 19.16.

Table 19-14: Metallurgical Recoveries and Concentrate Grade Used to Estimate Mineral Reserves

	Units	Hypogene		Mesogene		Supergene
Metallurgical Recovery Cu	%	92.7	%	88.1	%	79.5	%
Metallurgical Recovery Au	%	89.0	%	83.1	%	68.8	%
Concentrate Grade	Cu%	27.0	%	27.6	%	58.1	%

Table 19-15: Metal Prices and Other Parameters Used to Estimate Mineral Reserve

Parameter	Units		Value
Concentrate Moisture Content	%		8.0	%
Exchange Rate	Can$ : US$		0.88
Copper Price	US$/lb	$	1.45
Gold Price	US$/oz	$	475.00
Concentrate Transport	Can$/t (wet)	$	54.00
Concentrate Shipping	US$/t (wet)	$	40.00
Concentrate Treatment	US$/t (wet)	$	80.00
Cu Refining	US$/lb	$	0.08
Payable Copper	%		96.6	%
Payable Gold	%		97.1	%
Copper Price Participation*	US cents/lb		2.5
Arsenic**	US$/dmt	$	2.50
Mercury***	US$/dmt	$	2.00

Notes: The contribution to ore value from silver is very minor and has been excluded from the estimation process.
*Copper price participation = 10% above a threshold price of US$1.20 per lb.
**Arsenic penalty= US $2.50 per dmt of concentrate for each 0.1% over 0.2%
***Mercury penalty = US $2.00 per dmt of concentrate for each 10 ppm over 10 ppm.

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Table 19-16: Description of Fields in the Combined Resource Model

	Coordinates of block centroids inside the following limits:
X, Y, Z	X from 3000 to 4150, Y from 1650 to 2200, Z from 4600 to 5750.
cu_comb	Merged block grades for copper, gold and silver calculated as the weighted averages of the resource
au_comb	block grades and the halo block grades.
ag_comb
As and Hg	As (g/t) and Hg (ppb) grades.
	Net Smelter Return at a copper price of US$1.45/lb and a gold price of US$475/oz. Silver values were
NSR	ignored in the NSR calculation.
	Block bulk density, estimated from the relative rock type proportions and based on densities of 2.57
Density	t/m3 for the Supergene, and 2.61 t/m3 for Hypogene, Mesogene and Waste.
Class	Resource model classification (1= Measured, 2=Indicated).
pct_ore	Classified resource volume percent for each block.
	Integer rock codes for Hypogene (2001), Mesogene (2002), Supergene (2003), Picrite (998), Dilution
Rock Type	Halo (997) and Waste (996). Blocks straddling the resource model and the halo model were assigned
(RKTYPE)	the resource model rock type.
	Domain codes for Unconsolidated surface material (100), Picrite (101), Hangingwall Fault (102), Fault
GTDOM	Impacted Zone (103), Hypogene(104), Mesogene (105), Supergene (106), Halo (107) and Waste
	(108).

The processes involved in preparing the resource model resulted in minor (<0.1%) differences between the total contents of the original resource model and the Combined Resource Model used for reserve estimation. The differences were mainly due to the different methods used by GEMS and Datamine software for calculating the contents of blocks that intersected the boundaries between the various geological and geotechnical domains, including the surface topography. A series of visual checks were also carried out to ensure that no significant errors in the distribution of mineralization and the classification of the resource had occurred.

19.10.3 Initial Selection of Mining Outlines

AMC undertook an initial review of the mineral distribution in the resource model to determine those areas most likely to support a block cave mining operation. A plan view of the model highlighting the different zones of mineralization is shown in Figure 19-3. The Hanging Wall Lenses are isolated from the Main Zone and would require separate underground infrastructure, should they be mined. These lenses are considered to be uneconomic at foreseeable long term metal prices and have been excluded from the reserve. The Pit Protection Pillar has also been excluded from the reserve because it contains slightly lower grade mineralization than the bulk of the Main Zone, and because it could not be easily be recovered as part of the East Block.

The block caving method requires the orebody to be undercut, enabling the ore to collapse and fragment (cave) into underlying drawpoints, where it is extracted. The shape and size of the area to be undercut (the footprint) was determined by comparing a nominated cost of establishing each drawpoint ($300,000) with the recoverable NSR Value of the overlying column of ore, after deducting mining, processing and all other on-site costs. A nominal cost of $15/tonne was used for this purpose. The footprint was designed to encompass those drawpoints with a positive net value, and where necessary, a number of sub-economic drawpoints to ensure that the undercut area was large enough to cave (determined by geotechnical investigations as being an area with a minimum span in any direction of approximately 95m).

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Figure 19-3: Plan View of Preliminary Resource Block Model

The heights of the caved ore columns were determined by applying a shut-off value to the ore recovered from drawpoints. Once the NSR Value of the ore being drawn reached less than C$15 (the shut-off value) the drawpoint was closed.

The analysis determined that the Main Zone mineralization could be mined as three cave areas with Block 1 (B1) and Block 2 (B2) separated by a low-grade pillar. B1 and B2 would be extracted from drawpoints on the same level, whilst Block 3 (B3) would be extracted from a lower elevation. Figure 19-4 shows a view (looking north) of the three caving areas relative to the open pit.

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Figure 19-4: Block Cave Outline Relative to the Existing Open Pit (Looking North)

19.10.4 Estimating Block Cave Production

Two separate software programmes were used to estimate the mineral reserve. PC-BC³ software was used to prepare an initial production schedule, and the three-dimensional particle flow code Cave-Sim⁴ was used to model the progressive mixing of various mineralized and unmineralized materials within the cave as mining progressed.

To assess the sensitivity of the reserve estimate to changes in key assumptions used to model the movement and recovery of ore from the cave, three cases were modeled and described as the Expected, the Best Credible and Worse Credible cases.

19.10.4.1 Modeling Using PC-BC

Key inputs to the PC-BC modelling process include the estimated shape of the draw cones, the percentage of fines, the mixing horizon and the sequence of extraction. The inputs are described below and summarised in Table 19-17.

Draw cones are treated in PC-BC as static shapes, which remain constant over the life of the drawpoint. Three draw cone profiles were modeled, one for each case. Narrow cone profiles, particularly in B1 and B2, were generated to reflect the fine fragmentation expected at the drawpoints and the expected limited interaction immediately above the extraction level.

³ Block cave reserve estimation software, a product of Gemcom Software International Inc.
⁴ A software code developed to help analyse the performance of different caving projects by direct simulation of granular flows of individual particles.

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The percentage of fines material within PC-BC relates to rock sizes having a cube length less than 20 cm. The default value is 30% fines, but this input was altered for each case to reflect expected differences in fragmentation.

An extraction sequence starting in the SW corner of B2 and progressing to the NW 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 45Ú to the strike of the orebody.

The scheduling method used aimed to close off all the drawpoints at about the same time. To achieve this, the far eastern drawpoints in B1 were “trimmed” to a maximum Height of Draw of 350m.

Table 19-17: Key Input Parameters to PC-BC Models

	Best Credible Case	Expected Case	Worst Credible Case
Cone Shape B1 & B2	Medium	Fine	Fine
Cone Shape B3	Medium	Medium	Fine
Percentage Fines	40%	60%	80%
PC-BC Mixing Horizon	60m	75m	100m
Values Common to all Cases
Maximum Height of Draw (HOH)			350m
Shut-off Value (NSR value)			C$15
Discount Rate			8%
Draw Point Construction Cost			C$300,000
Maximum Drawbell Construction Rate per Month			8
Maximum Undercutting Rate (m2/month)			1,352
Scheduling Interval			Quarterly
*Planned indicative draw-down rate
Months 1 to 3			50 mm/d
Months 4 to 6			100 mm/d
Months 6 to 9			200 mm/d
Month 10 onwards			400 mm/d

*Actual rate varies due to cone shape and production requirements

The results from the cases run using PC-BC are summarised in Table 19-18.

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Table 19-18: Summary Results from PC-BC Modelling

	PC-BC Schedules
	Best Credible Case	Expected Case	Worst Credible Case
B1 ‘000t	19,630	18,358	13,693
B2 ‘000t	16,407	15,339	11,534
B3 ‘000t	10,719	9,850	9,149
TOTAL ‘000t	46,736	43,532	34,261
B1 Cu%	1.12	1.14	1.14
B2 Cu%	0.90	0.89	0.88
B3 Cu%	0.89	0.89	0.89
TOTAL Cu%	0.99	0.99	0.98
B1 Au g/t	0.74	0.75	0.74
B2 Au g/t	0.73	0.72	0.71
B3 Au g/t	0.69	0.68	0.67
TOTAL Au g/t	0.72	0.73	0.72
NSR Value B1+B2 (C$ million)	1,145	1,075	797
NSR Value B3 (C$ million)	323	297	272
TOTAL NSR Value (C$ million)	1,468	1,372	1,070
AVERAGE NSR Value C$/t	31.42	31.52	31.22

19.10.4.2 Modeling Using Cave-Sim

As Cave-Sim does not schedule an optimum extraction rate from each draw point, a schedule developed using PC-BC was therefore used to control the rate of extraction, except that in the Expected and Best Credible cases, extraction from each draw point was allowed to continued until 90% of the scheduled tonnage had been drawn. Extraction then continue only as long as the drawpoint grade (NSR Value) remained above the shut-off value. In the Worst Credible case, an 80% cap was applied before grade control took over.

Eight different particle flow domains were assigned for numerical analysis. These domains were selected on the basis of their particle flow parameters, principally the D50 size fraction. Within each domain, three sets of primary (P) and secondary (S) fragmentation estimates were applied for each case. The fragmentation estimates used are shown in Table 19-19 and Table 19-20.

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Table 19-19: Key D50 Fragmentation Inputs to Cave-Sim (Main Ore Types)

	Supergene		Hypogene		Hypogene		Mesogene
			B2		B3		B1
	P	S	P	S	P	S	P	S
Best Credible
Block Volume (m3)	0.050	0.020	0.200	0.030	0.400	0.200	0.080	0.015
Cube Edge Length (m)	0.368	0.271	0.585	0.311	0.737	0.585	0.431	0.247
Equiv. Spherical dia. (m)	0.457	0.337	0.726	0.386	0.914	0.726	0.535	0.306
Expected Values
Block Volume (m3)	0.020	0.005	0.020	0.010	0.080	0.030	0.080	0.030
Cube Edge Length (m)	0.271	0.171	0.271	0.215	0.431	0.311	0.431	0.311
Equiv. Spherical dia. (m)	0.337	0.212	0.337	0.267	0.535	0.386	0.535	0.386
Worst Credible
Block Volume (m3)	0.015	0.001	0.015	0.005	0.080	0.015	0.080	0.015
Cube Edge Length (m)	0.247	0.100	0.247	0.171	0.431	0.247	0.431	0.247
Equiv. Spherical dia. (m)	0.306	0.124	0.306	0.212	0.535	0.306	0.535	0.306

Table 19-20: Key D50 Fragmentation Inputs to Cave-Sim for Other Rock Types

	Halo	Waste	Picrite	FIZ*
	S	S	S	S
Best Credible
Block Volume (m3)	0.025	0.025	0.001	0.001
Cube Edge Length (m)	0.292	0.292	0.100	0.100
Equiv. Spherical dia. (m)	0.363	0.363	0.124	0.124
Expected Values
Block Volume (m3)	0.010	0.010	0.001	0.001
Cube Edge Length (m)	0.215	0.215	0.100	0.100
Equiv. Spherical dia. (m)	0.267	0.267	0.124	0.124
Worst Credible
Block Volume (m3)	0.005	0.005	0.001	0.001
Cube Edge Length (m)	0.171	0.171	0.079	0.100
Equiv. Spherical dia. (m)	0.212	0.212	0.098	0.124

*Fault Impacted Zone (FIZ)

The Cave-Sim modelling process tracked the movement of various material types through the cave and reported the production of copper, gold, silver, mercury, arsenic grades and NSR Value by quarter for each key rock type (Hypogene, Mesogene, Supergene, Halo, Waste and Picrite). The results from the three cases run are summarised in Table 19-21.

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Table 19-21: Summary Results* from Cave-Sim Modelling

	Cave-Sim Schedules (C$15/t Shut-off)
	Best Credible Case	Expected Case	Worst Credible Case
B1 ‘000t	20,780	20,176	18,065
B2 ‘000t	16,631	15,971	14,511
B3 ‘000t	9,764	9,447	8,819
Total ‘000t	47,175	45,594	41,396
B1 Cu%	1.06	1.08	1.02
B2 Cu%	0.86	0.88	0.84
B3 Cu%	0.87	0.88	0.83
Average Cu%	0.95	0.97	0.92
B1 Au g/t	0.70	0.71	0.68
B2 Au g/t	0.71	0.72	0.69
B3 Au g/t	0.69	0.69	0.65
Average Au g/t	0.70	0.71	0.68
NSR Value B1 (C$ million)	636	630	538
NSR Value B2 (C$ million)	495	488	422
NSR Value B3 (C$ million)	289	284	250
Total NSR Value (C$ million)	1,421	1,402	1,210
Average NSR Value C$/t	30.11	30.76	29.22

* Silver grades not shown

19.10.4.3 Comparisons between PC-BC and Cave-Sim Results

Both PC-BC and Cave-Sim provide estimates of the total recovered tonnages and grades, enabling a comparison to be made between the output from the two methods. In the Expected Case, Cave-Sim modelling indicates marginally higher levels of dilution than PC-BC, resulting in higher tonnages (+5%) and lower grades (-2% to -3%) and the total recovered NSR Value (a proxy for recovered metal) using the Cave-Sim method is approximately 2% greater than using PC-BC.

19.10.5 Adjustment for Ore Losses

In AMC’s opinion mining losses will occur in addition to those simulated by Cave-Sim and PC-BC software, in particular the following:

• Over the life of the project there will be deviations from the draw control rules used in the simulations that will result in waste entering the draw earlier than expected.

• Other than in the best ground conditions, it is expected that some drawpoints will become damaged beyond economic repair before reaching their shut-off grade.

• Towards the end of the mine life there will be some drawpoints that have not reached their shut-off grade, but the total number of drawpoints in this condition will be insufficient to justify continued operation of the mine.

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To account for these effects and other miscellaneous losses, the output from the Cave-Sim Expected Case was adjusted by prematurely stopping some drawpoints in each block before they reached their shut-off grade. Minor adjustments were then made to the rate of draw from each remaining drawpoint to achieve the scheduled mine production rate. The adjustments resulted in an overall reduction of approximately 4% in the tonnage output from the Cave-Sim Expected Case.

Ore recovered from drawpoints after adjusting for losses is shown in Table 19-22, and is summarised in Table 19-23.

Table 19-22: Results from Cave-Sim Expected Case After Losses*

B1 ‘000t	19,450
B2 ‘000t	15,674
B3 ‘000t	8,834
Total ‘000t	43,958
B1 Cu%	1.09
B2 Cu%	0.88
B3 Cu%	0.90
Average Cu%	0.98
B1 Au g/t	0.728
B2 Au g/t	0.73
B3 Au g/t	0.71
Average Au g/t	0.72
NSR B1 (C$ million)	615
NSR B2 (C$ million)	481
NSR B3 (C$ million)	272
Total NSR (C$ million)	1,368
Average NSR C$/t	31.13

* Silver grades not shown

Table 19-23: Summary of Ore Recovered from Drawpoints

	Tonnes	Cu	Au	Ag	NSR Value
	(Mt)	(%)	(g/t)	(g/t)	(C$/t)
Ore recovered from drawpoints	44.0	0.98	0.72	2.27	31.13

19.10.6 Development Ore

A small quantity of ore, not included in the block cave modeling work, will be mined during development. The estimated quantity of development ore is shown in Table 19-24.

Table 19-24: Development Ore

	Tonnes	Cu	Au	Ag
	(Mt)	(%)	g/t	g/t
Development ore	0.4	0.69	0.62	1.76

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19.10.7 Total Ore Mined

The total quantity of ore mined from development and from the block cave drawpoints is shown in Table 19-25.

Table 19-25: Ore Recovered from Development and Drawpoints

	Tonnes	Cu	Au	Ag	NSR Value
	(Mt)	(%)	(g/t)	(g/t)	(C$/t)
Ore recovered from development and drawpoints	44.4	0.98	0.72	2.27	31.13

19.10.8 Mineral Reserve Classification

The Measured and Indicated Resources lying within the vertical projection of the footprint that are expected to cave are shown in Table 19-26. No Inferred Resources are included in the outline.

Table 19-26: Mineral Resources Contained within the Vertical Projection of the Cave Footprint

	Tonnes	Cu	Au	Ag
	(Mt)	(%)	g/t	g/t
Measured Mineral Resource	31.4	1.22	0.89	2.75
Indicated Mineral Resource	8.9	0.83	0.73	2.50
Inferred Mineral Resource	-	-	-	-
Total	40.2*	1.13	0.86	2.69

* Totals do not equal the sum of the components because of rounding adjustments

Although a large volume of the material encompassed by the projected footprint has a Measured Resource classification, a Probable Reserve classification has been applied to the reserve because of the uncertainty associated with estimating material movement within the cave, and the absence of any historical actual versus forecast reconciliation at Afton to provide guidance. Also, a large quantity of Indicated Resource and unclassified material mixes with the Measured Resource within the cave. As it is not possible to mine the Measured Resource separately from this material, the effect is to lower the classification of the total reserve.

19.10.9 Mineral Reserve Estimate

The Mineral Reserve estimate for the New Afton Project, announced April 2, 2007, is presented in Table 19-27. The Reserve has been estimated using the resource block model prepared by Scott Wilson RPA, the contents of which form the basis of the Mineral Resource Estimate summarised in Table 19.12. The entire mineral reserve is included in the Mineral Resource Estimate. The reserve has been reported using the CIM Definition Standards on Mineral Resources and Mineral Reserves Definitions and Guidelines.

Table 19-27: Mineral Reserve Estimate¹

	Tonnes	Cu	Au	Ag	NSR Value
	(Mt)	(%)	(g/t)	(g/t)	(C$/t)
Probable Mineral Reserve	44.4	0.98	0.72	2.27	31.13

1      Estimated at US$1.45 Cu/lb and, US$475/oz Au using a cut-off NSR Value of C$15/t of ore, metallurgical recoveries, treatment charges and other parameters shown in Tables 19-14 and 19-15.

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20. Other Relevant Data and Information (This section was written by John

Shillabeer, P.Eng., Hatch).

20.1 Execution Plan

The mine development and operation until July 2011 will be managed by NGD directly and implemented by a mining contractor, Cementation Canada, (Cementation). It has been assumed for the Feasibility Study that NGD will award the EPCM contract for the surface facilities and infrastructure to an experienced EPCM contractor.

The Owner: As construction is completed and start-up and commissioning begin, the permanent operations staff will assume control of the mill and surface infrastructure. NGD will establish a project organization to direct the project; review and approve design; approve purchase decisions and contract awards; manage the treasury and operate the accounts payable department; liaise with governments; obtain the necessary permits and establish the permanent operating organization. In addition to these project-wide functions, NGD will establish a team to manage Cementation.

The EPCM Contractor: The EPCM contractor will be responsible for basic and detail design; specifications and procurement of equipment and materials; development and packaging of construction contracts; as well as the supervision of construction activities. The EPCM contractor will implement systems to monitor, control and report progress against project budget and schedule. The EPCM contract will be open book with bonus and penalty provisions for safety, schedule and cost. Reporting to the NGD Chief Operating Officer (“COO”), the EPCM project manager will be responsible for delivering the project as defined. This will make for a clear division of corporate accountabilities.

The Mining Contractor: Cementation will be responsible for developing the mine in accordance with the study mine plan. The mining contractor will operate the mine in the initial phase while continuing to develop it until the expansion part of the project is completed and production ramps up to 4 Mt/y. Cementation will implement systems to monitor, control and report progress against project budget and schedule ensuring that the desired throughput is achieved. Cementation will be required to provide safety, environmental, design engineering, procurement, site engineering, project schedule and cost controls, labour relations, general supervision and management for all the underground works including initial operations reporting to NGD’s COO. Cementation, will be retained on an open book, target price, risk sharing type of contract. This type of contract allows the flexibility for engineering design to continue during development, and shares risks between those most able to manage risk.

The NGD COO will focus on preparing the transition to operations, set up a system to manage change orders and discharge those project responsibilities which are within the Owner’s scope.

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The EPCM project manager will manage the development of basic and detail engineering, procurement of process equipment and materials, construction contracts formation and bidding, project procedures for managing the project which include and is not limited to engineering, procurement, project controls, contract administration, construction, QA/QC, Safety, health and environmental. The EPCM project manager will integrate consultants as required by the project and ensure they meet the projects expectations including review and approval of health, safety and environmental plans and procedures proposed by contractors. Enforcement will be first by contractors, backed up by the project manager. The EPCM contractor will complete procurement based on specifications developed by the engineers. For continuity, technical consultants who completed site investigations and initial designs described in this document will continue to advise the project.

In general, the preferred method of contracting for work not within Cementation’s scope, is to use fixed prices with sufficient detailed engineering completed at the time of bidding to permit the use of realistically priced hard money contracts. However, a limited amount of incentive based contracting may be used on critical path items where circumstances warrant. In order to ensure that project objectives are met, the project schedule, defined work scopes and budgets will be the EPCM team’s primary means of project execution control and the baseline against which change will be managed.

Changes to the overall scope of the project are not expected.

The NGD President will establish a formal delegation of authority with commitment limits for key members of the project team before the project is initiated. NGD’s COO will maintain contingency and escalation funds globally and use an established management procedure to manage transfers.

The project schedule is determined by the time required to develop the mine and the current long delivery times being experienced for SAG mills. In order to achieve commercial production by mid 2009 NGD has commenced work on the two long lead items identified in the FS, underground works and SAG mill delivery.

Cementation has mobilized to site and will commence underground works in April 2007. Initially this work will be carried out under authority of the existing exploration permit while the Chief Inspector of Mines is adjudicating the project approval permit.

NGD has placed orders for the main SAG mill components and motor for delivery in fourth quarter of 2008.

The underground development schedule is aggressive but achievable. There is modest float available in the underground development and construction schedule and hence that part of the project will be “schedule-driven”. The overall project construction strategy will be strongly influenced by:

Reaching the extraction level as soon as possible from the existing exploration development and disposing of waste rock in the open pit.

Establishing through ventilation via a fresh air raise as early as possible to remove a major constraint on the amount of development that can be achieved per day.

Using relatively small, immediately available underground development equipment to best effect until the 50 t capacity trucks and large jumbo are available for duty on the main ramp.

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Surface construction is not “schedule-driven” except at the outset when it is desirable to have the main access road, plant site earth works, buried services, permanent power and water supplies completed in 2007 to facilitate the mining development and to minimize costs.

A summary of the project development schedule is included as Figure 20-1

20.2 Other information

Hatch is not aware of any additional information or explanations beyond those contained in this report required to make this document understandable and not misleading.

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21. Interpretations and Conclusions

21.1 General (This section was written by John Shillabeer P.Eng., Hatch)

The site investigations; geological, geotechnical, metallurgical and survey data; engineering designs and environmental studies; cost estimates and economic modelling conducted to date meet the standards expected of a detailed feasibility study. The feasibility study is sufficiently detailed and the assumptions so defined, that NGD is now able to make decisions about committing the Project to construction.

21.2 Project Economics (This section was written by John Shillabeer P.Eng., Hatch)

Hatch can offer no comment on the future of metal prices, exchange rates or costs, as these are all outside the control of the project participants. These parameters have significant impacts on project economics. Nonetheless, subject to the relevant qualifications assumptions and exclusions set out in this report and the satisfactory completion of the additional work recommended in section 22, the project would appear to be profitable at the base date of this report.

Using the 3 year historical average metal prices and exchange rate listed in section 25.8 and no escalation, the project achieves an after tax rate of return equal to 10.4% and payback in 6.3 years.

21.3 Geology (This section was written by David Rennie, P. Eng., Scott Wilson RPA)

Scott Wilson RPA has completed a mineral resource estimate for the New Afton deposit. At a C$10/t gross dollar value cut-off, the present estimate comprises Measured and Indicated Resources totalling 65.7 Mt grading 1.02% Cu, 0.77 g/t Au, and 2.59 g/t Ag.

The estimate was carried out using diamond drill results collected and compiled by NGD personnel. In Scott Wilson RPA’s opinion, the sampling, assaying, and assay QA/QC protocols used at New Afton are appropriate for the deposit type and have been carried out to a standard suitable for estimation of mineral resources. The assay database has been properly compiled, is relatively free of errors, and is suitable for use in mineral resource estimation.

The updated estimate has a higher proportion of Measured to Indicated Resources from earlier estimates due to the increased density of the drilling, which has improved the overall confidence level of the estimate. The present drill section spacing is a nominal 40 m, which is approximately one half of what it was prior to the start of the 2005 in-fill drilling program. With completion of the in-fill drilling, the general form and location of the deposit did not change significantly from earlier estimates. This demonstrates, in Scott Wilson RPA’s opinion, that there is good geological continuity from section to section and the location of the zone can be predicted from the geological model with a fair degree of certainty. For this reason, Scott Wilson RPA considers all mineralization contained within the interpreted geological boundaries (i. e., the wireframes) to be in at least the Indicated category.

An additional resource body, termed the C Zone, has recently been discovered below the Main Zone ore body. The C Zone mineral resources at the C$10/t cut-off are estimated 7.94 Mt grading 0.96% Cu, 0.88 g/t Au, and 1.55 g/t Ag. The C Zone was not included in the Mineral Reserves estimate.

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Additional block models have been constructed for the dilution halo around the block cave volume, and for the Hg and As content of the mined volume. These models are preliminary in nature owing to a lack of sample information. Additional sampling and assaying is required to allow for more accurate estimates of Hg and As in the deposit.

21.4 Mining (This section was written by Mike Thomas, MAusIMM (CP)., AMC Consultants Pty Ltd.)

After consideration and analysis of a range of other mining methods, the block caving method was selected for the deposit. This method provides the best economic return, while at the same time providing the highest resource recovery and the lowest technical risk.

Both empirical and numerical assessments indicate that the orebody will cave naturally at the design undercut dimensions.

The deposit is highly fractured and the proportion of the cave rock mass that will pass 2 m³ (a standard industry guideline) will be close to 100% for all areas.

The orebody is capable of sustaining a production rate of 4 mtpa.

The block caving method is a bottom up method, which will enable mining to commence lower in the deposit in hypogene material and for most of the development to be located in more competent rock.

The Probable Mineral Reserve of the New Afton Project is 44.4 Mt grading 0.98% Cu, 0.72 g/t Au and 2.27 g/t Ag, based on the conditions stated in Section 19. The mineral reserve was classified as Probable Reserve because of the uncertainty associated with estimating movements of natural materials within the cave zone.

21.5 Mineral Processing (This section was written by Ken Major, P. Eng., Hatch)

The process design criteria and process flowsheet for NGD’s Project have been interpreted by Hatch from the results generated by the metallurgical test program that has been completed at SGS Lakefield located in Ontario.

The metallurgical prediction models have been interpreted from the test results using statistical and mathematical methods of data analysis.

For detailed interpretation of results refer to Section 18.

21.6 Environmental Permitting (This section was written by Rolf Schmitt, P.Geo, Rescan Environmental Services)

The New Afton Project Application for a Permit Approving the Mine Plan and Reclamation Program under the B.C. Mines Act, was accepted for government technical review and public comment in January 2007. Once approvals are obtained, authorization will be obtained to proceed with permitting. A number of government permits are required and will be expedited by the agencies through the lead role of the Chief Inspector of Mines under the Ministry of Energy, Mines and Petroleum Resources.

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The Project’s complex geology, glacial history and continental, semi-arid climate have given rise to a variety of landscapes. Much of the project area has been greatly altered by historic mining, mineral exploration, grazing, utility and transportation corridor activities such that only remnants of natural landscapes remain. Within the mining lease area, the dominant landscapes are now classified as brownfield and include level rock dumps, open pits, tailings storage, compacted borrow areas, and several man-made alkaline bodies of water. The New Afton project will occupy approximately 234 ha of the total 902.9 ha mining lease, of which about two-thirds of this land (152 ha) will occupy lands previously disturbed by mining and on-third (81 ha) will alter grazing lands. The post-closure reclamation program will add a net addition to reclaimed lands at the Afton site.

Baseline environmental studies focused on the area’s surface water and groundwater, wildlife, heritage, soil and acid rock drainage potential. Surface water and groundwater is naturally alkaline and shows a range of water quality and trace metal constituents. Hydrological and groundwater modeling determined that all surface and subsurface flow paths will be within the proposed footprint of the mine and be directed to Afton pit where waters will be naturally contained. Extensive hydrological analysis of the pre-Afton mine, current, and future site conditions have consistently concluded that there will be zero surface discharge from the site and that all groundwater within the mine footprint will flow towards and collect in the Afton pit and the underground mining void (after closure). Sixty-nine wildlife species and eighty-seven plant species were identified and environmental management plans are proposed to address the potential presence of species at risk and noxious weeds. The mine lease area does not contain any fish or fish habitat. A chance find recovery procedure will be implemented in the event any unknown sites of archaeological or heritage significance are encountered during construction. Soils mapping identified areas of soil suitable for salvage and use for reclamation, and provided information on natural agricultural capability and vegetation that will assist in establishing mine permit reclamation objectives. Finally, extensive metal leaching and acid rock drainage assessment and prediction studies determined that it is highly unlikely that the tailings facility will generate acid rock drainage.

Environmental management plans have been developed to guide the monitoring, mitigation and remediation of impacts that may occur during construction, mine operation and mine closure activities. Water quality monitoring, suppression of dust and the protection of wildlife are features of the environmental management plans. A comprehensive reclamation program is included wherein lands disturbed by mining will be restored to a pre-determined land use capable of supporting cattle grazing and wildlife.

Water conservation is paramount in the dry interior grasslands of the Thompson Plateau. The project design will ensure that there is no surface water discharge from the site, and that all groundwater within the mining footprint will flow towards and collect in Afton Pit. The waste rock from the underground operations will be minimal and will be deposited in the bottom of the existing Afton Pit prior to flooding at closure. Nearly all waste rock has been determined as non-acid generating, therefore the final deposition of waste rock in the Afton pit, and subsequent flooding will ensure that the potential for any acid generation is highly unlikely. Tailings have also been determined to be non-acid generating and will be deposited in the southeast of the mining lease where all surface and groundwater will be directed to Afton pit, and the surface of the tailings facility will be revegetated at closure to support grazing and wildlife use.

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22. Recommendations

The costs of implementing all of the recommendations listed below are included in the applicable project capital cost or operating cost estimates listed in tables 25-7,25-8 and 25-16. There is once exception which is the recommendation to continue exploration drilling to find additional resources. The project capital estimate does not include exploration spending. The costs of assessing project opportunities (section 22.8) are included in project estimates

22.1 Geology (This section was written by David Rennie, P. Eng., Scott Wilson RPA)

In Scott Wilson RPA’s opinion, the sampling and geology work at New Afton is being carried out in an appropriate fashion to common industry standards. However, there are a few areas which could benefit from additional information or analysis in order to improve the mineral resource estimate. Scott Wilson RPA makes the following recommendations:

• Additional geological interpretation should be carried out to try and resolve the controls on mineralization, particularly in the areas immediately outside of the present resource body.

• Routine analyses for As and Hg should be added to the assay protocols in order to provide the basis for block modeling of these elements. A re-assay program of existing sample rejects is being undertaken to improve the database for As and Hg.

• Additional geological interpretation work should be undertaken to determine the controls to the As and Hg mineralization.

• Exploration diamond drilling should be continued to find additional mineral resources.

22.2 Mining (This section was written by Mike Thomas, MAusIMM (CP)., AMC Consultants Pty Ltd.)

Assessment of the potential mudrush hazard has been based on limited information. In particular, the information available for determining the potential for fines generation within the cave zone is inadequate. AMC recommends that rock property tests be carried out, in particular slake durability tests, for the specific rock types representative of those expected within the cave.

Further assessment of subsidence predictions is recommended, which should include refinements to the stress analysis model to include major structural features and a degree of calibration through mining of an additional crosscut parallel to the existing exploration crosscut. The behaviour of the open pit to date should be back analysed as an additional step in calibrating the model.

During operations, establishment of comprehensive subsidence monitoring systems and rigorous management procedures are recommended to provide the earliest possible verification of the predictions and reasons for any deviations.

It is proposed that the extraction level and undercut be formed using modern drill and blast techniques. However, there may be advantages to using a roadheader for this work and it is recommended that this method be investigated.

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Further geotechnical investigations are recommended for the proposed ventilation shaft locations. These investigations may indicate that alternative methods, sizes and types of shaft to those proposed in the Feasibility Study may be more suitable.

There is potential to simplify and reduce the cost of the proposed decline conveyor transfer point arrangements. Further engineering and geotechnical investigations relating to the decline alignment and to the transfer point arrangements are recommended prior to finalising the detailed design of these facilities.

The Feasibility Study design specifies the use of a fire resistant belt, a fire detection system along the length of the decline and sprinklers at the head and tail ends. The adequacy of this approach to mitigating fire risk may need to be further assessed.

The accuracy of mineral reserve estimates in block caving operations is influenced to a large degree by the ability to predict the draw characteristics of the caved rock mass. Draw characteristics are sensitive to the size distribution of the caved material, with fine fragmentation generally leading to lower recovery. To mitigate the risk AMC recommend that a program of further drilling and fragmentation analysis be carried out during the mine development stage to improve the accuracy of the fragmentation assessment and to finalise the block cave design and operating strategies.

There is a possibility that a lack of caving experience on site could lead to poor designs and poor understanding of the factors necessary for operating a successful block caving mine, resulting in adverse economic and safety impacts. AMC recommend that this be addressed by recruiting suitably experienced personnel and using external expertise to fill any gaps.

Additional geological and geotechnical data is required for the final design of the B1 and B2 undercut and extraction levels and for other specific areas where underground excavations such as the crusher chamber will be required. A drilling program designed to increase the geological and geotechnical understanding of conditions in theses areas is recommended. The program should be carried out when the access decline reaches a suitable position.

AMC recommend that additional stress measurements be undertaken at depth when suitable sites become available. This, together with structural mapping and modelling during development should identify any areas requiring additional support.

It is envisaged that ongoing site investigation drilling will determine the ground conditions for specific excavations (e.g., ventilation shafts, conveyor decline), and that stability analyses will be conducted from the results of these investigations during a detailed engineering stage.

22.3 Metallurgy and Processing (This section was written by Ken Major, P. Eng., Hatch)

Although the primary processes of grinding and flotation have been effectively represented in the metallurgical test program and in the development of the metallurgical projections, additional work should be considered in the areas of secondary processes (concentrate filtration) and the dilution effects of block caving.

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The concentrate filtration circuit has been developed in conjunction with the equipment vendors based on operating experience at other copper mills. To develop sufficient concentrate from lab tests would require additional sample for testing.

The dilution work that has been completed for the New Afton project indicated that the waste rock did not negatively influence the magnitude of the value minerals recovered in the flotation process using the bulk underground samples. This program could be extended to the variability test samples to increase the data points available for the analysis.

22.4 Surface Geotechnical (This section was written by Monte Christie, P.E., Vector Engineering Inc.)

Prior to detailed design, a Phase 2 geotechnical field investigation should be performed to verify material properties used in the design assumptions for the TSF and Pothook facilities. In particular, the existing waste dump should be drilled and sampled to verify amount of compressibility beneath the TSF dam and stored tailings. Testing should also be performed on durability of subgrade materials subjected to process solution. The final design should also verify settlement of stored tailings and design preliminary closure plans. The Pothook dam and foundation should be designed. Surface water diversions around both the TSF and Pothook facilities should be finalized. And specific monitoring instrumentation should be recommended, including settlement markers and pore pressure piezometers.

22.5 Environmental Permitting (This section was written by Rolf Schmitt, P.Geo., Rescan Environmental Services Ltd.)

NGD is recommended to continue on-site environmental monitoring in order to meet regulatory permitting conditions and to support the present environmental data base which informs the effectiveness of the environmental management plans. The current monitoring program should be modified for construction and operational phases while continuing to conform to regulatory requirements.

NGD has already prepared environmental management plans and an emergency response plan. These should be elaborated and implemented during construction.

• Enhance Environmental Management plans to guide the impact management, monitoring, remediation and reporting of environmental impacts and values during the development, operation, closure and post- closure of the New Afton Mine.

• Develop an environmental monitoring plan for emissions, surface water and groundwater, wildlife and vegetation in accordance with conditions established by the provincial regulatory agencies. The results of monitoring should be regularly reviewed and adjustments made to the sampling protocol if required. It is anticipated that the frequency and scope of monitoring activities will progressively reduce during the phases of mine operations, through closure to post-closure.

• Continue the kinetic testing of selected pyrite zone rock samples for a period of time until such excavated material is brought to surface during mining, to better inform long-term physical placement and handling of the material.

• Ensure sufficient high quality soil is identified, salvaged, stockpiled and protected for progressive and final reclamation and revegetation of the mine site.

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• Implement the Chance-find-recovery procedure for sites and objects of archaeological and heritage interest during construction.

• Ensure all mine employees are familiar with the Company’s environmental policy , objectives and procedures for the minesite.

22.6 Groundwater (This section was written by Andrew Holmes P Eng., Piteau Associates)

NGD is recommended to install and operate the proposed seepage interception wells early in the construction phase. This will allow for further calibration of the numerical groundwater flow model, and will intercept seepage from the existing Afton tailings facility towards the open pit during construction.

22.7 Construction (This section was written by John Shillabeer, P Eng., Hatch)

NGD is recommended to negotiate a management contract with an experienced EPCM Contractor and to continue advancing preparations for mine development and SAG mill fabrication in order to protect the integrity of the Project schedule.

22.8 Project Opportunities (This section was written by John Shillabeer, P Eng., Hatch)

The project team identified a number of areas with the potential to enhance project outcomes. These opportunities were not included in the project scope because more investigative work will be required before they can be relied upon.

22.8.1 “C“ Zone Resource

Additional drilling to delimit and upgrade the classification of the C Zone resource to Measured and Indicated categories has the potential to add additional life and operational flexibility to the project.

22.8.2 Early Ramp up to 4 Mt/y

The project schedule allows two years of initial operations before the underground crusher and conveyor are completed and production begins to increase to 4Mt/y. This interval is proposed in order to:

• Minimize the time and capital cost exposure before generating revenue.

• Build up site-specific cave management expertise.

If initiation and early operation of the cave is achieved according to plan, management may consider the two year interval to be excessive and decide to bring forward the ramp up. However, careful analysis will be required of the commercial and risk issues, together with the accelerated plan for early development and equipping of the crusher-belt conveyor system.

22.8.3 Delete Temporary Ore Crusher

Present plans require a contractor to re-handle and crush all run of mine ore delivered by truck to the surface before it is fed into the SAG mill. There will be no size control underground, except for the size of the LHD bucket. The ore is predicted to break relatively finely. More detailed observation and analysis of rock mass characteristics during mine development, together with initial experience of cave behavior may show that the crusher will not be required. Possibly a grizzly over the feeder will be sufficient. Savings could be of the order of $1/t ore for 3.2 Mt of ore produced in the first two years.

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22.8.4 Optimize Cave Draw Schedule

There is scope to further optimize the cave draw schedule to mine higher grade material earlier and to delay drawing ore with higher arsenic values. A program should be carried out to optimize the draw point development rate and the rates of draw for each draw point.

The distribution of arsenic (and mercury) is not as well understood as those of copper and gold, because the data base of arsenic assays is not as large as for copper and gold. NGD is carrying out further As and Hg assays to develop a better understanding of the As distribution, when it will be appropriate to re-examine the schedule.

22.8.5 Two Tier Electricity Tariff

The cost of electrical energy consumed by the project was estimated with reference to BC Hydro tariff Schedule 1823, effective July 2006. This tariff provides for a flat rate of 2.852 cents per kWh to be charged to new customers for twelve consecutive billing periods. The project capital and operating cost estimates use the flat rate throughout the project life.

After twelve billing periods, the tariff provides for:

Tier 1: 2.569 cents per kWh for all kWh up to 90% of the previous consumption (calculated annually)

Tier 2: 5.400 cents per kWh for all kWh above 90% of the previous consumption (calculated annually)

This rate structure is both an opportunity and a challenge. Clearly, if operations were entirely constant , there would be an incentive to reduce consumption and capture the savings of the tier 1 tariff. However, operations will not be constant because of the two stages of project development and subsequently because of fluctuating development and operations in the mine. The details of the tariff provide methods to establish a baseline for new or expanded operations. Therefore NGD can plan to minimize its exposure to Tier 2 charges and capture the savings from the Tier 1 rate.

22.8.6 Delay Pit Debris Stabilization

Saturated debris which has accumulated in the bottom of the Afton Pit must be stabilized to minimize risks to underground operations. Stabilization methods are described in Section 25.1.10. In the presently proposed caving sequence, Block 2 will be mined in advance of Block 1. Block 2 cave is not predicted to intersect the pit bottom, whereas Block 1 cave is predicted to intersect the pit floor (and the debris). Additional studies of rock mass behavior, detailed mapping and interpretation of structures may allow some schedule flexibility to be introduced. For example, it may be possible to safely delay the completion of debris stabilization until just before the Block 1 cave begins. The cost of stabilization would be deferred and probably taken as a sustaining capital cost. Some operational (patrol) costs would be avoided.

22.8.7 Automation of Underground Operations

Block caving lends itself to a high level of automation and remote control. It is recommended that detailed discussions be entered into with suppliers in order to maximize the benefits of semi- or fully– automated mucking on the extraction and ore transfer levels.

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The mine communications and control plan and cost estimate provides for a high capacity fiber optic backbone to accommodate future remote control, monitoring and automation.

22.8.8 Sub Level Caving Under The Block Cave

The resource tapers off into a keel below the block cave zones. Extraction of this keel may be possible using sub level caving after completing the block cave above. This has not been studied in detail in the feasibility study because the of the low impact on the net present value of the project as it would occur at the end of the life of the mine.

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23. References (This Section was Written by John Shillabeer, P.Eng., Hatch)

(1)	Behre Dolbear & Company, Ltd.; January 2004; Mineral Resource Estimate For the Afton Copper/Gold Project, Kamloops, B. C.; internal report to DRC Resources Corporation; 160 pp.
(2)	Currie, James A.; February 2004; Advanced Scoping Study For the Afton Copper/Gold Project, Kamloops, B.C.; internal report to DRC Resources Corporation; 187 pp.
(3)	Deutsch, C. V. and Journel, A. G.; 1998; GSLIB Geostatistical Software Library and User’s Guide, 2nd Edition; Oxford University Press, Inc.; 369 pp.
(4)	Konst, R.; April 2006; New Afton Project, 2005-2006 Drilling Program, Sample Preparation and Analytical Quality Control Report; internal report to New Gold Inc.; 17 pp.
(5)	New Gold Inc. Press Releases and Financial Statements; from website, www.newgoldinc.com.
(6)	Stanley, Clifford R.; 1994; Geology of the Pothook Alkalic Copper-Gold Porphyry Deposit, Afton Mining Camp, British Columbia (92I/9, 10); Mineral Deposit Research Unit (MDRU), University of British Columbia, Contribution #035, Geological Fieldwork, Paper 1994-1; 10 pp.
(7)	Whiteaker, R. J., and Hall, R.;September, 2006; Assessment Report For the New Afton Project, Underground Development, Diamond Drilling, Geochemistry, and Geological Work; New Gold report filed with the BC government; 33 pp.
(8)	World Climate website, www.worldclimate.com.
(9)	AMC’s Afton Project Mining Study Report
(10)	McDonald and Harbough, 1988. “A Modular Finite-Difference Ground-Water Flow Model.” Book 6, Modelling Techniques, USGS Publication.
(11)	Piteau Associates Engineering Ltd., 2006. “Permit Level Hydrogeological Assessment for Tailings Storage Facility and Block Cave.” Report prepared for New Gold Inc., November, 49 p.
(12)	Rescan Environmental Services Ltd., 2006. “New Afton Copper-Gold Project – Application for a Mines Act Permit Approving the Mine Plan and Reclamation Program.” November.
(13)	SGS Metallurgical Services – report on metallurgical test program to come
(14)	“Metallurgical Test Report”, Project No. KRTS 20231, Knelson Research & Technology Center, August 23, 2006
(15)	“Gravity Modeling Report”, Project No. KRTS 20231-1 Rev.1, Knelson Research & Technology Center, September 5, 2006
(16)	“An Investigation into Arsenic Speciation in Six Copper Concentrate Samples”, Project 11236-001 Report 1, SGS Lakefield Research Ltd., October 2, 2006
(17)	“Flocculant Screening, Gravity Sedimentation, Pulp Rheology, Pressure Filtration and Vacuum Filtration Studies”, Pocock Industrial Inc., May 2006
(18)	“Deportment of Gold in Composites of Ore Styles from New Afton Mine”, Report 06/03, AMTEL Ltd, September 11, 2006

References
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(19)	“Petrographic Examinations of Rock Samples from The New Afton Project”, Report 060366, Vancouver Petrographics Ltd., June 16, 2006
(20)	“Mineralogical Examination of Samples from the U/G Bulk Sampling Program at The New Afton Property”, Report 060466, Vancouver Petrographics Ltd, July 11, 2006
(21)	B.C. Ministry of Forests, 2004. Forest Practices Code of B.C. Act, Regulations and Guidelines.
(22)	BC Ministry of Environment. 2002. Stormwater Planning: A Guidebook for British Columbia.
(23)	Canadian Dam Association, 1999. CDA Canadian Dam Safety Guidelines. 118 p.
(24)	Coulson, C.H. 1991. Manual of Operational Hydrology in British Columbia. Ministry of Environment, Water Management Division, Hydrology Section, B.C., Canada.
(25)	Coulson, C.H. and W. Obedkoff. 1998. British Columbia Streamflow Inventory. Water Inventory Section, Resource Inventory Branch, Province of British Columbia Ministry of Environment, Lands, and Parks. 49 pp.
(26)	Hogg, W. and Carr, D. 1985. Rainfall Frequency Atlas for Canada. CIAACC. Centre de Ressources en Impacts et Adaptations au Climat et a ses Changements, Canada. http://www.criacc.qc.ca/villes/idf_e.html
(27)	Preliminary Geotechnical Investigations of The Proposed Plant Sites: Brian Nachtigal; BGC Engineering Inc; July 2006
(28)	Preliminary Geotechnical Field Investigation Report For The Proposed Mill Plant Site At The New Afton Copper And Gold Mine; Anon; Vector Engineering Inc; October 2006
(29)	Pre Feasibility Study Of Afton Mine Water Supply For New Gold Kamloops BC: Peter Coxon; Urban systems Ltd; December 2006
(30)	Tailings Pond Peak Flows: Brian Nachtigal; BCC Engineering Inc.; April 2006
(31)	CP Rail/Arrow quote dated June 12, 2006.
(32)	HVC quotation dated January 4, 2007.
(33)	Ocean Freight - Simpson Spence & Young Estimate, December 2006
(34)	Arrow update re cost of loading trucks (email to Kevin Ross from Joe Mather, January 19, 2007)
(35)	Afton Operating Corporation. 1996. Mine Closure Plan.
(36)	Afton Operating Corporation. 1998. 1997 Reclamation Report. Self-published report.
(37)	Afton Operating Corporation. 1999. 1998 Reclamation Report, Reclamation Permit M-112. Prepared for the Ministry of Energy and Mines by Afton Operating Corporation. March, 1999.
(38)	Afton Operating Corporation. 2001. 2000 Reclamation Report, Reclamation Permit M-112. Prepared for the Ministry of Energy and Mines by Afton Operating Corporation. March, 2001.
(39)	Afton Operating Corporation. 2004. 2003 Reclamation Report, Reclamation Permit M-112. Prepared for the Ministry of Energy and Mines by Afton Operating Corporation. March, 2004.

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(40)	B.C. Ministry of Environment, Lands, and Parks (MELP) and B.C. Ministry of Forests (MoF). 1998. Field manual for describing terrestrial ecosystems. Land Management Handbook No. 25. Victoria, B.C BC Weed Control Act 2001 http://www.qp.gov.bc.ca/statreg/reg/W/66_85.htm
(41)	BC Ministry of Environment. 1978. Kamloops Lake Reservoir and Storage Compilation.
(42)	BC Ministry of Environment. 1981. Climate Capability Classification for Agriculture in British Columbia, APD Technical Paper 4. Victoria, BC..
(43)	BC Research. 1992. Urban Runoff Quality Control Guidelines. Prepared for Fisheries and Oceans Canada and BC Ministry of Environment by BC Research. 1992.
(44)	BC Weed Control Act 2001 http://www.qp.gov.bc.ca/statreg/reg/W/66_85.htm
(45)	British Columbia Ministry of Environment. 2005. Wetlands in B.C. http://wlapwww.gov.bc.ca/wld/
(46)	wetlands.html. Accessed October 2005.
(47)	British Columbia Ministry of Water, Land and Air Protection (BC MWLAP). 2004. Best Management Practices for Amphibians and Reptiles in Urban and Rural Environments in British Columbia. B.C.
	MWLAP Ecosystem Standards and Planning Biodiversity Branch. Nanaimo, B.C.
(48)	Cavanagh, N., R.N. Nordin, L.W. Pommen and L.G. Swain, 1998. Guidelines for Designing and Implementing a Water Quality Monitoring Program in British Columbia. BC Ministry of Environment, Lands and Parks. Resource Inventory Commission, Forest Renewal British Columbia.
(49)	Clark, J. 2003. Invasive Plant Prevention Guidelines. Centre for Invasive Plant Management, Bozeman, Montana.
(50)	Environment Canada. 2002. Metal Mining Guidance Document for Aquatic Environmental Effects Monitoring. Environment Canada, Ottawa. June 2002.
(51)	Environment Canada. 2005. National Air Pollution Surveillance (NAPS) Network – Annual Data Summary for 2003. Environmental Technology Advancement Directorate, Environmental Protection Service. Totonto, Ontario.
(52)	Environment Canada. 2005. The New Metal Mining Effluent Regulations. Environment Canada, Ottawa. Accessed at: http://www.ec.gc.ca/nopp/docs/regs/mmer/en/index.cfm.
(53)	Environment Canada. Greenhouse Gas Division. 2005. Canada’s Greenhouse Gas Inventory: 1990– 2003. Ottawa, ON.
(54)	Meteorological Services of Canada, accessed in August, 2006. (http://www.climate.weatheroffice.ec.gc.ca/climate_normals/index_e.html)
(55)	Ministry of Energy, Mines and Petroleum Resources. 2003. Health, Safety and Reclamation Code for Mines in British Columbia.
(56)	New Gold Inc. 2006. New Afton Mines Permit Application Terms of Reference. V 8.1. August, 2006.
(57)	New Gold Inc. 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, Volume 1 – Mine Permit Application, Volume 2 – Appendices.

References
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(58)	Province of British Columbia. Mines Act RSBC 1996, Chapter 293. Queens Printer, Victoria, British Columbia, Canada.
(59)	Rescan. 2006. Interpretation of ML/ARD Static Testwork - New Afton Project. August 2006.
(60)	Ames, S.A. 2000. Impacts of Land Use on Carbon Storage and Assimilation Rates. PhD Thesis. Resource Management and Environmental Sciences. UBC, Vanccouver, BC.
(61)	Ames, S.E., Lavkulich, L.M . 1999. Predicting the role of land use on carbon storage and assimilation rates. World Resource Review. Vol. 11, no. 1, pp. 30-46.
(62)	Lapsa, Melissa, Trends in Renewable Energy: Focus on Copper, Global Market Trends Conference 2006, Sept. 2006
(63)	Goodyear, Chip, Chief Executive Officer, BHP Billiton, Where To From Here? Merrill Lynch Global Metals, Mining and Steel Conference, May 2006,
	www.	bhpbilliton.com/bbContentRepository/Presentations/060327CWGML2006ConfMiami.pdf
(64)	Barry, Sean, High prices, rising fuel costs crippling copper fabricators; substitution, weaker demand taking a tol,lAmerican Metal Market, Sept 2005, http://www.findarticles.com/p/articles/mi_m3MKT/is_36-1_113/ai_n15634884
(65)	Xstrata completes acquisition of Falconbridge Ltd. under rules for remaining shares, Oct 13, 2006, http://www.cbc.ca/cp/business/061013/b101330.html
(66)	Organisation for Economic Co-operation and Development, Economic Projections, http://www.oecd.org/topicstatsportal/0,2647,en_2825_32066506_1_1_1_1_1,00.html
(67)	Mbendi, World: Mining - Copper Mining, http://www.mbendi.co.za/indy/ming/cppr/p0005.htm
(68)	Global InfoMine, http://www.infomine.com/commodities/copper.asp
(69)	RusMet, Metallurgical Industry of Russia, World copper mine production grows 34% from 1996- 2005: ICSG, Aug. 2006, http://www.rusmet.net/news.php?act=by_id&news_id=50635
(70)	Investment Moats, Copper: The Metal with the PH.D, http://www.investmentmoats.com/?p=93
(71)	Kitco Base Metals, http://www.kitcometals.com/
(72)	Stablum, Anna, Reuters: Copper, zinc and aluminium prolong metals rally, Oct, 2006, http://za.today.reuters.com/news/newsArticle.aspx?type=businessNews&storyID=2006-10- 18T103641Z_01_BAN822733_RTRIDST_0_OZABS-MARKETS-METALS-20061018.XML
(73)	Copper Development Assn., www.copper.org
(74)	Westhouse, Copper Market Overview, December 2005
(75)	ALTA Metallurgical Services, Copper SX/EW – Any Rivals in Sight?, February 2002
(76)	ICSG PRESS RELEASE: New Edition of ‘Directory of Copper Mines and Plants’, June 2006,
	www.	icsg.org
(77)	Aur Resources Inc., Investors Presentation, Registration at the “Bolsa Off Shore”, 2006

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(78)	Australian Chamber of Commerce and Industry, ISSUES PAPER, OECD Businesses Expect Strong Growth in 2006 but Slowdown in 2007, July 2006
(79)	Cumerio, General information on copper market, August 2006, www.cumerio.com
(80)	Edelstein, Daniel L., US Geological Survey, COPPER
(81)	Storer, Robyn, Serious Wake-up, World Mining Stocks, March 2006
(82)	Barrick Gold Corporation 2005 Annual Report, http://www.barrick.com/Theme/Barrick/files/docs_annual/2005_AR_en_rev2.pdf
(83)	Gold Market Review and Outlook, Prepared by GFMS Ltd on behalf of the CBOT®, June, 2006
(84)	New Afton Feasibility Study, Financial Model and Valuation, Draft 21 Dec, 2006: file name Financial Model R4.xls
(85)	BGC Engineering, Inc., 2006, “Drilling and Instrumentation in the Proposed Tailings Storage Facility (TSF)”, New Afton Project Feasibility Study, May 28, 2006.
(86)	Bishop, A., and Bjerrum, L., 1960. “The relevance of the Triaxial Test to the Solution of Stability Problems,” Research Conf. on Shear Strength of Cohesive Soils, ASCE, University of Colorado, pp.
	437-501.
(87)	Cedergren, H.R., 1988, “Seepage, Drainage, and Flow Nets”, 3rd edition, John Wiley & Sons. Inc., New York, NY.
(88)	Coduto, D.P., 2002, “Geotechnical Engineering Principles and Practice”, Prentice-Hall of India, New Delhi.
(89)	Darcy, H., 1856, “Les Fountaines Publiques de la villa de Dijon”.
(90)	Department of the Army U.S. Army corps of Engineers, 1986, “Engineering and Design Seepage Analysis and Control for Dams”, EM 1110-2-1901, Washington. D.C.
(91)	Holtz, R.D. and Kovacs, W.D., 1981, “An Introduction to Geotechnical Engineering”, Prentice Hall, Englewood Cliffs, New Jersey.
(92)	Koerner, R.M., 1998, “Designing with Geosynthetics”, fourth edition, Prentice Hall, Englewood Cliffs, New Jersey.
(93)	Johnson, S. 1975, “Analysis and Design relating to Embankments”, Analysis and Design in Geotechnical Engineering, ASCE, Vol.2, pp.1-48.
(94)	Lowe, J, 1967. “Stability Analysis of Embankments”, Journal of Soil Mechanics and Foundations Division, ASCE, Vol. 93, No. SM4, pp. 1-34.
(95)	Logan, J.M.,2003, “Iron Mask Project, Kamloops Area: British Columbia”, Geological Survey, Geological Fieldwork 2002, Paper 2003-1.
(96)	Ministry of Energy, Mines, and Petroleum Resources, Minefile Number 092INE023.
(97)	Rocscience, Inc, 2003. “SLIDE-5.0; A 2D Limit Equilibrium Slope Stability for Soil and Rock Slopes”, Toronto, Canada.

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(98)	Rodgers, Gary, 2006, Geologic Survey of Canada -Personal Communication.
(99)	Seed, H.B., and Idriss, I.M., (1982), “Ground motions and Soil Liquefaction During Earthquakes”, Earthquake Engineering Research Institute, University of California at Berkeley.
(100)	SLIDE, 2003, (Rocscience, 2003) “2D Limit Equilibrium Slope Stability for Soil and Rock Slopes”, Version 5.027 dated October 13, 2006, Rocscience, Inc., Toronto, Canada.
(101)	Spencer, E., 1967, “A Method of Analysis of the Stability of Embankments Assuming Parallel Inter- Slice Forces”, Geotechnique, Vol. 17, No1.
(102)	Spencer, E., 1967. “A Method of Analysis of the Stability of Embankments Assuming Parallel Inter- Slice Forces”, Geotechnique, Vol. 17, No. 1.
(103)	Tokimatsu, K., and Seed, H.B. (1987), “Evaluation of Settlements in Sands Due to Earthquake Shaking”, Journal of Geotechnical Engineering, v. 113, No. 8, August, pp. 861-878.
(104)	Youd, L.T., Idriss, I.M., Andrus, R.D., Arango, I., Castro, G., Christian, J.T., Dobry, R., Finn, W.D.L., Harder, L.F. Jr., Hynes, M.E., Ishihara, K., Koester, J.P., Liao, S.S., Marcuson, W.F. III, Martin, G.R., Mitchell, J.K., Moriwaki, Y., Power, M.S., Robertson, P.K., Seed, R.B., and Stokoe, K.H. II (2001), “Liquefaction Resistance of Soils”, Summary Report From the 1996 NCEER and 1998 NCEER/NSF
	Workshops on Evaluation of Liquefaction Resistance of Soils, Journal of Geotechnical and Geoenvironmental Engineering, October, pp. 817 – 833.
(105)	Youd, L.T., Hansen, C.M., and Bartlett, S.F. (2002), “Revised Multilinear Regression Equations for Prediction of Lateral Spread Displacement”, Journal of Geotechnical and Geoenvironmental Engineering, December, 1007- p. 1017.
(106)	Bruce Geotechnical Consultants Inc., 2006a. “Drilling and Instrumentation in the Proposed Tailings Storage Facility.” Memorandum prepared for Hatch Associates, regarding the New Afton Project Feasibility Study. May 28, 6p.
(107)	Bruce Geotechnical Consultants Inc., 2006b. “Summary of Hydrogeology Field Work and Results.” Memorandum prepared for Hatch Associates, regarding the New Afton Project Feasibility Study, July 21, 9p.
(108)	Klohn Leonoff Consultants, 1977. “Report on Piezometer Installation, Tailings Dam – Afton Mine.” January 28, 7p.
(109)	McDonald and Harbough, 1988. “A Modular Finite-Difference Ground-Water Flow Model.” Book 6, Modelling Techniques, USGS Publication.
(110)	Piteau Associates Engineering Ltd., 2006. “Permit Level Hydrogeological Assessment for Tailings Storage Facility and Block Cave.” Report prepared for New Gold Inc., November, 49p.
(111)	Rescan Environmental Services Ltd., 2006. “New Afton Copper-Gold Project – Application for a Mines Act Permit Approving the Mine Plan and Reclamation Program.” November.

References
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24. Date and Signature

24.1 John Shillabeer, P.Eng., Hatch Ltd

Qualified Personnel Letters
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24.2 Ken Major, P.Eng., Hatch Ltd.

Qualified Personnel Letters
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24.3 David Rennie, P.Eng., Scott Wilson RPA

Qualified Personnel Letters
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Andrew Holmes, P.Eng., Piteau Associates Engineering Ltd.

Qualified Personnel Letters
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24.4 Monte Christie, P.E., Vector Engineering Inc.

Qualified Personnel Letters
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24.5 Rolf Schmitt, P.Geo., Rescan Environmental Services Ltd.

Qualified Personnel Letters
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24.6 Mike Thomas, MAusIMM (CP), AMC Consultants Pty. Ltd.

Qualified Personnel Letters
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24.7 Mike Struthers, C. Eng, MAusIMM, MIMMM, AMC Consultants (UK) Ltd.

Qualified Personnel Letters
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25. Additional Requirements for Technical Reports on Development Properties and Production Properties

25.1 Mining Operations

This section was written by Mike Thomas and Mike Struthers of AMC Consultants Pty Ltd. Mr. Struthers wrote sections 25.1.2 through 25.1.9 inclusive and Mr. Thomas wrote the remainder, (except that section 25.1.10 was extracted from a report provided by MEG. )

In this Section, references to direction, North, South, East, West etc relate to the Mine Grid, in which North is orientated 50Ú west of the UTM grid. The origin and bearing of the mine grid relative to the UTM grid is as follows:

Mine Grid Origin - 674673.041E, 5610753.638N.

Bearing - 310Ú.

References to elevations relate to a mine datum set at 5,000 m below UTM datum.

25.1.1 Mining Method

The block caving method proposed for the deposit was selected after consideration of the following mining methods:

Non-Caving Methods

Open pit mining

Sub-level 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)

Sub-level caving (SLC)

Block caving

Key conclusions from the various studies carried out during the selection process are as follows.

A ‘Whittle’ pit optimisation study carried out to assess the potential for open pit mining eliminated this option as uneconomic due to the high stripping ratio required.

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An initial geotechnical assessment of the rock mass quality indicated that open stoping spans in the order of 15-20 m would be possible in hypogene mineralisation, but that stable stoping dimensions in mesogene mineralisation would generally be less than 15 m, and more typically 10 m, which is a practical minimum for open stoping. The assessment also indicated that open stoping in the supergene ore would not be practicable. Because only small open stopes would be possible, operating costs and consequently cut-off grades would be high compared to open stoping mines operating in better ground conditions. At the higher cut-off grades required, the deposit tends to break down into discontinuous zones of mineralisation. As a result, these methods would be unable to economically recover large portions of the deposit.

The SLC method requires mining to commence at the top of the deposit and progress in a top down sequence. If this method were to be used, mining would need to commence in areas of poorer rock quality associated with the supergene and mesogene ore types. These ore types also contain higher arsenic and mercury grades than the hypogene ore type.

The block caving method is a bottom up method, which will enable mining to commence lower in the deposit in hypogene material and for most of the development to be located in more competent rock.

Economic analysis and risk assessment of the SLC and block caving methods indicated that block caving would provide the best economic return, while at the same time providing the highest resource recovery and the lowest technical risk. The method selection studies also indicated that there was potential to mine some small remnant areas below the block cave panels by the SLC method later in the mine life.

As a result of the mining method selection studies, further studies were carried out to determine the optimum block caving arrangements, resulting in three block caving panels, Block 1 (B1), Block 2 (B2) and Block 3 (B3). Figure 25-1 shows the location of the blocks relative to the existing open pit

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Figure 25-1: Block Cave Outline Relative to the Existing Open Pit (Looking North)

25.1.2 Geotechnical Domains

After a series of investigations, the following geotechnical domains were defined:

• Supergene mineralisation

• Transitional (Mesogene) mineralisation

• Primary (Hypogene) mineralisation

• Fault-Impacted Zone

• Footwall (waste) rock

The Supergene, Mesogene and Hypogene geotechnical domains match the geological domains defined by the nature of the copper mineralization. The Fault Impacted Zone (FIZ) is a narrow (0-20 m wide) zone of extremely poor ground conditions immediately adjacent and parallel to the Hangingwall Fault.

In addition to the geotechnical domains, there is a picrite body to the south of the Hangingwall Fault in close proximity to the orebody for a short strike length.

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

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25.1.3 Structural Features

The most prominent fault sets impacting on mine design are the:

• Hangingwall Fault Zone

• Footwall fault

• North-east faults

• North-west 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 mineralisation, and the worst within the intensely altered FIZ, the Hangingwall Fault Zone, the adjacent picrite, and where major faults intersect the Supergene mineralisation.

25.1.4 Geotechnical data

The current project drill hole database comprises 207 holes with a total of 90,981m logged in the geology file, and 86,974m in the geotechnical file. Drilling coverage is good within and close to the mineralised zones, but decreases as the distance from the main mineralised areas increases. Specific drilling has provided geotechnical information in areas of key mine infrastructure. There is clear evidence of deterioration of core over time, in the weaker materials (e.g., weakest Supergene, some picrites).

All geotechnical logging by NGD has been based upon Bieniawski’s original 1976 version of his rockmass rating system, the RMR76. AMC reviewed the raw geotechnical data provided by NGD, and generated RMR76 values during post-processing.

Table 25-1 and Figure 25-2 summarise the basic geotechnical data for each of the domains.

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Table 25-1: Geotechnical Statistics by Domain

		RQD	F/m	Jt. Cond.	Strength Index	RMR76	Q Estimate
Hypogene	Mean	70	10.3	16	6	45	1.12
	Median	76	6.3	20	7	48	1.56
	Std Dev.	25	8.0	6	2	11	0.03
Mesogene	Mean	66	9.9	14	6	42	0.81
	Median	74	6.2	12	7	45	1.12
	Std. Dev.	28	8.3	7	2	12	0.03
Supergene	Mean	51	12.2	11	4	34	0.33
	Median	60	16.2	12	4	37	0.46
	Std. Dev.	32	7.2	6	2	12	0.03
FIZ	Mean	39	14.0	10	4	31	0.23
	Median	40	16.1	12	4	32	0.26
	Std. Dev.	32	6.8	6	2	12	0.03
Footwall	Mean	60	10.8	16	6	43	0.89
	Median	65	7.0	20	7	45	1.12
	Std. Dev.	24	6.7	6	2	10	0.02

To facilitate an empirical assessment of cavability, a conversion to Laubscher’s RMR (1990 version) was developed, based on re-logging of selected boreholes. This correlation is considered satisfactory for broad indications of behaviour, such as in a cavability assessment. The resulting best-fit correlation is as follows:

RMR90 = (RMR76 - 15) / 0.786

The correlation is reasonable above RMR76 = 45, but relatively poor below. This is suited to the main objective of the cavability assessment – the investigation of cavability in the more competent materials.

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Figure 25-2: Distribution of RMR76 by Domain (All Data)

Based on a limited programme of laboratory strength testing and field point load index testing, the recommended rockmass unconfined compressive strengths for design purposes are:

• Hypogene 32MPa

• Mesogene 22MPa

• Supergene 15MPa

The virgin rock stress was measured by overcoring of HI cells in Hypogene mineralisation exposed in the underground exploration crosscut. Four cells were successfully overcored, but poor biaxial tests occurred in all cases. Ultimately, results from tests 1 and 4 were used as shown in Table 25-2.

Table 25-2: Virgin Stress Measurements Results

																Vertical
Test	Sigma 1	Dip		Bearing		Sigma 2	Dip		Bearing		Sigma 3	Dip		Bearing		Stress
	(MPa)					(MPa)					(MPa)					(MPa)
1	17.3	27	o	197	o	11.0	36	o	309	o	7.5	42	o	079	o	10.8
4	18.5	08	o	183	o	8.9	64	o	291	o	7.1	24	o	089	o	8.8

Depth below surface = 290 m
Bearings in relation to UTM, and dips positive down. For Mine Grid add 50o to azimuth

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25.1.5 Cavability

The cavability of a deposit defines the ability of the orebody and overlying rockmass to cave freely and spontaneously, once undercut to a sufficient dimension. It is generally understood that any rockmass will cave, providing the dimensions of the undercut are sufficient. The cavability of the New Afton deposit has been assessed using two techniques; (i) empirical and (ii) numerical stress analysis.

25.1.5.1 Empirical Cavability Assessment

The standard empirical assessment for cavability is the methodology developed by Laubscher (1990, 2001) whereby a parameter, the Mining Rock Mass Rating (MRMR) derived from the RMR90 data (see Laubscher, 1990) is plotted on Laubscher’s 1990 Stability Graph together with the Hydraulic Radius (HR) of the block to be caved. The HR is defined as HR = area/perimeter of the undercut. Hence for example, a 100m x 100m undercut would equate to an HR = 25 m.

The hydraulic radii for the three caving blocks at New Afton are:

B1 and B2: HR = 34 m 
B3: HR = 31 m

The following adjustments were applied to RMR90 to obtain the MRMR:

Adjustments
Weathering             0 
Orientation          85% 
Stress                   100% 
Blasting                    0 
Total                    0.85

Table 25-3 details the distribution of RMR90 for each block and for the rockmass in the area of each undercut (Ucut). Figure 25-2 shows these data plotted on Laubscher’s (1990) Stability Chart.

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Table 25-3: Summarised RMR90 by Domain

Cave	RMR90						MRMR
	20	%	50	%	80	%	20	%	50	%	80	%
Block 1 Cave	33		47		59		28		40		50
Block 1 Ucut	43		48		55		37		41		47
Block 2 Cave	41		53		62		35		45		53
Block 2 Ucut	43		53		65		37		45		55
Block 3 Cave	39		53		65		34		45		55
Block 3 Ucut	33		52		59		28		44		50

In each case the lower levels of the rockmass (the undercut) are slightly more competent than the upper levels (represented as the ‘cave’). This is especially true in B1, where the upper levels of the cave include a significant proportion of weaker Supergene ore.

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 25-3), the margin from the indicated “non-caving” zone is small. The north-south span for B3 in particular is recognised as a potential issue.

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Figure 25-3: Caving Stability Chart (after Laubscher, 1990)

25.1.5.2 Numerical Cavability Assessment

The three-dimensional finite element code Abaqus was used to assess the development of caving.

The Abaqus stress analyses indicated that a minimum north-south span of 95m was required, and that caving would initiate at an HR of 30 m. This was based on Hypogene rock properties. Again, this provides for

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adequate cavability for B1 and B2, but marginal cavability for B3. The cavability for B1 would be expected to be at a reduced HR, given the high proportion of the slightly weaker Mesogene rockmass within the cave volume.

25.1.5.3 Cavability Conclusions

Both the empirical and numerical assessments indicate that B1 and B2 will cave at the designed hydraulic radius. Although the cavability results for B3 are marginal, the analyses nevertheless suggest B3 will cave. In both the empirical and numerical assessments the role of major structures has not been considered. Major faults will assist cave development, especially in B1 which appears to contain a higher proportion of second-order faults (e.g., the north-east faults).

The structural environment associated with B3 is not currently well defined – this should be a focus for future data collection and investigative programmes. B3 will be mined later in the mine life, providing time for such studies. The risk of cave performance in B3 being poorer than expected can be mitigated through the use of cave assistance measures, plus the option of cave pre-conditioning if required.

25.1.6 Fragmentation

An assessment of the size distribution of broken rock at the extraction level drawpoints has been carried out using the BCF software program. The program provides an indication of the primary fragmentation resulting from the initial caving processes and the secondary fragmentation that occurs when material higher in the orebody is subjected to comminution as it is drawn down towards the drawpoints.

The work indicates that the Afton deposit is highly fractured within all geotechnical domains and the proportion of the caved rockmass that will pass 2m³ (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% <2m³ is possible. The secondary fragmentation remains high with 98% <2m³. Some large blocks may therefore be expected to report to the draw points in the early stages of production within this domain.

Overall there is likely to be only a small percentage of low hang-ups and oversize rock (greater than 2m³) reporting to the draw points.

Additional drilling is recommended to obtain high-quality structural data for further fragmentation analysis in key areas.

25.1.7 Subsidence

The potential for subsidence above the New Afton deposit has been evaluated by both empirical and numerical methods. The numerical assessment used a combination of draw simulation software, Cave-Sim, and a non-linear stress analysis software, Abaqus. However the numerical model was uncalibrated and included a number of important simplifying assumptions.

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The empirical assessment of cave limits provides a preliminary guide to possible surface effects. In practice, cave and subsidence limits may vary significantly from this estimate, mainly through variations in the actual bulking factors experienced, and varying degrees of influence of major geological features on cave propagation. The assessment used a bulking factor of 15%, together with an assumed initial minimum cave break angle of 80^o, and close to surface, slightly flatter crater slope angles compared to the existing pit slopes. Figure 25-4 shows the results of the empirical assessment, in relation to surface infrastructure. An exclusion zone for permanent or difficult to replace surface infrastructure has been established with a nominal stand-off distance of 100m from the empirical cave limits. Some ground movement can be expected to occur within the exclusion zone. The figure also shows the cave limit developed using the numerical method.

AMC recommend that further numerical assessment be carried out, which should include refinements to the Abaqus model to include major structural features (eg HWF, Picrite) and a degree of calibration through mining of an additional crosscut parallel to the existing exploration crosscut. The behaviour of the open pit to date should be back-analysed as an additional step in calibrating the stress analysis model.

Subsidence effects will develop progressively as mining takes place, with the maximum effect apparent only at the end of the mine life. The risk of subsidence extending outside the exclusion zone increases as the quantity of ore removed from the mine increases. AMC recommend establishment of comprehensive subsidence monitoring systems and rigorous management procedures to provide the earliest possible warning that subsidence effects may be deviating from predictions, and the reasons for the deviations.

Figure 25-4: Subsidence Limits Based on Empirical Assessment Methods

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25.1.8 Stability and 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 future during a period of refinement of B3 layout and designs. The proportions of different ground conditions and the required support requirements have been considered, based upon statistical analyses of geotechnical data.

The undercutting and extraction level development sequences have been designed to reduce damage from mining induced stresses. The proposed support systems for the undercut and extraction levels include 24mm-diameter full-column grouted rockbolts, cable bolts, 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 proposed to protect drawpoint pillar noses.

For the majority of the other excavations, stability assessments have been based upon statistical analyses of ground conditions and empirical estimates of support requirements, using the Q-System of rockmass classification. Extensive use of full column grouted rock bolts, cable bolts and fibre reinforced shotcrete is proposed .

Stress analysis studies of the proposed crusher chamber demonstrated that there is no interaction expected with the cave zones, and that the extent of yield around the crusher chamber will be limited to local reentrants, brows etc. Conventional support estimates have been made for the crusher complex.

It is envisaged that ongoing site investigation drilling will determine the ground conditions for specific excavations (eg. ventilation shafts, conveyor decline) and that stability analyses will be conducted from the results of these investigations during a detailed engineering stage.

25.1.9 Mudrush Potential

The potential for mudrush conditions to develop is a feature of many block cave mines. An assessment made of the mudrush issue at New Afton concluded that there is potential for a mudrush hazard to develop during the life of the mine, due to the potential for the generation of fine material in the cave mixing with water entering the cave. There is also the potential for ingress of wet mud and fines from the Afton pit, if this material is not effectively stabilised.

A series of engineering measures and operating procedures are proposed, which focus on minimising the entry of water into the cave and minimise the accumulation of fines within the cave, by developing strict draw control strategies.

25.1.10 Pit Dewatering and Debris Stabilization

This section is a condensed version of a report provided by MEG.

After open pit operations ceased in 1987, the bottom of the pit became infilled with slide debris and water. The accumulation of slide debris and water in the bottom of the open pit presents an unacceptable hazard to the proposed caving operation. Operators of block caving mines have recognized the potential for a sudden,

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uncontrolled inrush of water and debris to the mine workings. This phenomenon has been referred to as Mud Rush .

It is estimated that there are approximately 2.6 Mm³ of supernatant water and 0.6 Mm³ m of rock and soil debris in the open pit. These estimates are based on the results of bathymetric surveys , and a comparison of current and historical airphoto surveys. The water and debris are both estimated to be about 50 m deep. Slides contributing to the accumulation of debris have been recorded at several locations around the open pit (Figure 25-5). No direct measurements of the soil thickness in the base of the pit nor the depth of water have been made.

Figure 25-5: General View of Slope Conditions at Afton Mine Pit

The total volume of water and debris is increasing at the rate of about 0.15 M m³ per year. Of this amount some 2-5,000 m³ is rock and soil. The majority of the debris is likely to have originated from the sedimentary rocks on the north side of the pit and from the weathered intrusive rocks of the south side. Based on this historical data, it is understood that failures on the north side of the open pit have been more extensive and have contributed more material to the total debris volume than the slopes on the south side.

25.1.10.1 Preliminary debris characterization

While the total volumes of water and slide debris that constitute the hazard to the proposed block caving operation are known with reasonable accuracy, the characteristics and spatial variation of the debris materials is not known. Observation and samples show that the slide materials of sedimentary origin (from the north slopes) tend to be finer and soil-like, while materials from the south slopes are expected to be coarser, ranging from sand up to boulder sizes. Hence, it is expected that the material inside the pit would comprise a mixture of particle sizes ranging from clay through to large boulders.

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A program of surface sampling was conducted on the north and south slopes of the open pit, above the standing water level. The maximum particle size sampled was limited to 40 mm. Gradation analyses of selected samples show that the materials are well sorted.

Grading analyses were also performed on the surface samples after they had been submerged in water for a period of about 48 hours to simulate underwater deposition. Gradation analyses on the submerged samples show an increase in the total fines content of up to 55%, as a result of de-aggregation of the coarser fraction when immersed in water.

Atterberg limit tests on the fine portion of the submerged samples (material sieved on #40 mesh) indicate that the liquid and plastic limits are about 45% and 20%, respectively.

The Atterberg limits of a soil are important in terms of the soils potential behaviour. Fine material with a water content equal to, or greater, than its liquid limit behaves like a fluid. On the other hand, fine material with a water content equal to or lower that its plastic limit essentially behaves as a solid. This difference in response is the key to the Mud Rush mechanism and constitutes the basis for the definition of the mitigation measures presented herein.

Based on a visual inspection of the slopes in the open pit, it was estimated that the recovered samples represent about 80-90% of the true gradation of the slide debris. Hence, 10-20% of the slide material is expected to have a particle size greater than 40 mm.

The natural water content and permeability of the in-place debris material are not available and have been estimated based on certain assumptions. For the natural water content, it was assumed that the underwater slide debris is fully saturated. The range of water contents was selected as between 30% and 35% by weight.

The representative range of soil permeability for the slide debris was estimated to be in the range 10^-4 to 10^-6 m/s with an average of 10^-5 m/s. The estimates were based on the particle size gradation curves and experience with similar types of soils.

25.1.10.2 Debris stabilization

According to Butcher et al. (2000) four elements must be present for a mud rush to occur:

• Potential mud forming materials, i.e., a large amount of fine material;

• Water;

• A disturbing mechanism that could trigger the movement of the fine material and water (e.g., subsidence within the cave); and

• Entry points through which the mud and water mixture can enter the mine workings (e.g., cracks, fractures, the cave itself).

All four elements appear to be present at New Afton and there is concern regarding the potential for a mud rush phenomenon to occur.

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NGD has focussed the remedial measures on reducing the availability of free water and on controlling the characteristics of the in-place slide debris material. The debris stabilization scheme will be based upon the following control measures:

25.1.10.3 Controlling the Availability of Free Water

• The supernatant water is to be removed by pumping. A submersible pump will be suspended in the water and gradually lowered as the water is drawn down. The pump will deliver water to a tank and pump station on the shore and from there the water will be pumped through a surface pipeline to Pothook Pit. The capacity of Pothook Pit will be increased to 2.9 Mm³ by constructing an earth fill dam adjacent to it. The rate of drawdown of the water in the open pit will be limited, to ensure that pore pressures in the pit slopes have time to come to equilibrium, so avoiding additional instabilities and/or slope movements. One year has been allowed for removal of the surface free water.

• A bookkeeping system for water balance in the open pit will be maintained throughout the duration of pumping and for as long as necessary after the free water has been removed.

• The pit pumping system will be retained after the supernatant water has been removed.

• During operations a diversion system, sump and pumping system remote from the open pit will be established. The diversion system will be set up to avoid further entry of surface runoff into the pit.

25.1.10.4 Controlling the In-Situ Characteristics of the Slide Debris Material

• An east-west access road will be constructed on the exposed debris surface once the supernatant water has been removed. The road will begin from the east end of the pit. The access road will be constructed using locally available materials (waste rock from pit operation) reinforced with geomaterials.

• A series of dewatering wells will be drilled from the proposed access road and will completely penetrate the full thickness of the slide debris in the base of the open pit. Samples recovered during the drilling of the wells will be used to characterize the soil types forming the slide mass. The distribution of these materials along the access road will also permit the lateral and vertical material variability to be assessed.
  Pump tests will also be performed in the wells to determine the in situ hydraulic conductivity. This information will be used subsequently to design the final arrangement for the dewatering system.

• Up to 40 wells are estimated to be required. Each well will have a separate submersible pump. The dewatering system will deliver to a main sump from which water will be pumped to Pothook Pit.
  Pumping will continue until the moisture content of the fine debris has been reduced approximately to the liquid limit. Progress and the condition of the debris will be monitored throughout the dewatering process.

• Further dewatering and consolidation will be required beyond the liquid limit to dewater the fine materials to approximately midway between the liquid limit and the plastic limit so that mudflow can not occur. This will be achieved by either vacuum pumping using the existing wells or by mechanical compaction. The choice will be made after the initial wells have been drilled and the debris mass has been defined.

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• In terms of timing for the dewatering operation, the mine is not considered to be at risk during initial development and not until the cave, or cracks propagated by the cave, communicate with the pit water and debris. Caving from Block B2 is not projected to communicate with the bottom of the pit. The control measures will be completed (except for ongoing maintenance pumping) and verified before caving begins in Block B1.

25.1.11 Mine Layout and Access

Because the deposit is bounded on the southern side by the Hangingwall Fault, which is anticipated to be difficult to mine through, permanent access and underground infrastructure has been designed on the north (footwall) side of the deposit.

It is proposed to drive an access decline from the base of the existing exploration decline to provide early access to the undercut and extraction levels. The decline will be driven at 5.5m x 5.5m at a gradient of 1:6 (16.7%) down to the B1 & B2 extraction level and 1:7 (14.3%) below this, down to the B3 extraction level.

A new ramp (conveyor decline) will be developed from a portal located close to the proposed mill site at an elevation of 5,674m. The conveyor decline has been designed at a gradient of 1:6 (16.7%) with four straight legs, varying in length from 700m to 1,400m. Conveyor transfer stations have been designed at the head and tail end of each leg. The cross sectional dimensions, at 5.5m wide by 6.5m high, provides adequate clearance for a loaded 50t capacity truck and the conveyor structure which will be suspended from the roof of the decline.

Lateral accesses to the undercut and extraction levels, crusher, workshops and other infrastructure facilities have been designed at 5.5m x 5.5m to accommodate 50t capacity trucks and high capacity scoops. Roadways will be formed with imported aggregate, graded, rolled and compacted to form a low resistance running surface. Side drains will be formed for proper drainage.

The feasibility study mine layout is shown in Figures 25-6 to 25-14.

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Figure 25-6: Composite Plan of Mine Workings

Figure 25-7: Mine Plan Overlayed by Aerial Photograph

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Figure 25-8: Plan of B1 and B2 Undercut on 5085 Level

Figure 25-9: Plan of B1 and B2 Extraction Level on 5070 Level

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Figure 25-10: Plan of Ore Transfer Level and Crusher on 5055 Level

Figure 25-11: Ventilation and Drainage Level on 5040 Level

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Figure 25-12: Block 3 Undercut Level on 4065 Level

Figure 25-13: Block 3 Extraction Level on 4950 Level

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Figure 25-14 Block 3 Ventilation and Drainage Drive on 4930 Level

25.1.12 Block Cave Design

Analysis carried out to establish the optimum elevation of the undercuts for each block resulted in the establishment of the B1 & B2 undercut at a nominal elevation of 5,085 m and the B3 undercut at 4,965 m.

The undercuts will be mined from crosscuts driven at nominal centreline spacings 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, which 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 reliability.

Extraction level crosscuts will be accessed from an access drive on the north side of the deposit. An offset herringbone layout is proposed, with the access crosscuts, (after application of surface support) being suitable for the use of either diesel or electric scoops with a 6 – 7 t capacity. This layout enables the scoop to travel bucket first towards the draw points and to enter the draw points without changing direction. The scoop is always on the outby side of the active drawpoint, avoiding the need to reverse past the drawpoint.

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Ore will be trammed back to short 2.4 m diameter ore passes connected to the production crosscut. Only one scoop 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 scoops, or with development and ground support equipment that will need to use the access drive. To allow for the long life required, ground support standards on the extraction level will be high with extensive use of fibrecrete, rockbolts and cable bolts. Concrete roadways will be constructed in the extraction level crosscuts and drawpoints. Figures 25-15 and 25-16 show the proposed design of the undercut and extraction levels.

Figure 25-15: Undercut Design

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Figure 25-16: Plan View of Extraction Level and Drawpoint Design

25.1.13 Mine Production Rate

A staged production build up is proposed, initially to an annualized rate of 1.6 Mtpa for a nominal two-year period (“Phase 1”), increasing to a maximum of 4.0 Mtpa (“Phase 2”). The staged build up is designed to provide time to develop the best operating strategies and practices for the mine.

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25.1.14 Ore and Waste Transport

A trade-off study assessed the merits of using either shaft or conveyor haulage for ore transfer to surface. The study was based on conceptual underground mine layouts for two alternative ore handling systems delivering to a surface stockpile. The conveyor haulage option was adopted for the feasibility study for the following reasons:

• The conveyor option had the lowest capital and net present cost.

• No fatal flaws or risks were identified which cannot be adequately managed.

• Concerns relating to conflicts between conveyor operation and maintenance activities, and the use of the ramp as the main access were felt to be manageable.

The proposed ore handling system includes a 150 to 200 t capacity crusher dump pocket, designed to allow two-sided tipping by scoops and single side tipping by 50 t capacity trucks. The pocket will be installed above a 2.4 m wide apron feeder. The feeder will control the feed rate to a 2,000 mm x 1,500mm nominal size single toggle jaw crusher capable of handling a maximum 1,500 mm feed size. A fixed rock breaker installed at the crusher feed hopper will minimise production delays due to crusher bridges.

The crushed material will discharge to a pocket and bin of approximately 200 t 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.

During initial development of the mine and during Phase 1, all ore and waste will be trucked to the surface. 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 surface via the conveyor ramp will be placed on top of the future subsidence area. Total rock movement during Phase 1 will peak at approximately 5,500t per day.

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 waste transfer to surface. However, provision has also been made in the design for waste to be batched through the conveyor system if this proves to be a more economic option.

25.1.15 Mine Ventilation

During production, the primary ventilation circuit will utilise an exhaust system with main surface fans installed on three raisebored shafts. Three horizontally mounted axial flow fans are proposed to meet the required duty.

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The main intake airways will be the conveyor decline, the exploration/access decline and two raisebored shafts. The total required airflow through the mine during both Phase 1 and Phase 2 has been estimated at 420 m³/s. Exhaust raises will be raisebored at 3.5m diameter and will each have an airflow of approximately 140 m³/s. The final number and size of main exhaust raises may change pending geotechnical advice resulting from the proposed drilling program. Various secondary and booster fan arrangements 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. Dust extraction systems will be installed on orepasses, at conveyor transfer points and the 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 whilst 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 at Afton, approximately 23% of days 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 marginally above freezing.

25.1.16 Mine Infrastructure

To construct the crusher chamber, conveyor transfer bays and the other excavations required to accommodate major facilities, access drives will be driven into the proposed excavation, then enlarged in a systematic manner to form the finished chamber. Ground support in the form of rockbolts, cablebolts and fibrecrete will be installed as the chambers are developed.

A main underground workshop will be constructed adjacent to the extraction level to provide a maintenance facility of the mobile equipment fleet. An underground fuel bay will be provided adjacent to a dedicated return airway and fitted with fire suppression equipment. 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. An underground lunch room and office will be constructed on the main extraction level.

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.

It is not intended to reticulate compressed air throughout the 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 13.8 kV via a ring main system.

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

25.1.17 Emergency Egress

The project construction schedule has been designed such that 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. During Phase 1, the main access will be via the conveyor decline with secondary access to the open pit via the access and exploration declines. Wherever practical, two independent access points will be provided to all working areas.

At the commencement of Phase 2, caving will cause access through part of the exploration decline to be lost. Prior to this occurring, a service raise, 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.

It is proposed to utilise a standard Alimak system equipped with an eight-person car for the emergency hoisting system. The system will include an independent power supply for use in the event of a major power failure. Portable refuge chambers suitable for 15 persons have been proposed for use in areas where a permanent second means of egress cannot be provided, or where one is required during construction. The main lunch room will also be designed to act as a refuge chamber.

25.1.18 Production Schedule

A 24 month period is required from commencement of the project to firing of the first drawbell. Initial ore from development, undercutting and production drawpoints will be stockpiled in a temporary ROM stockpile area adjacent to the mill. Milling operations have been scheduled to commence in month 29, when the mine production rate will be sufficient to enable the mill to operate at its Phase 1 production rate (1.6 Mtpa). It is proposed that mining and milling will continue at this rate for a two year period before increasing to the Phase 2 rate of 4.0 Mtpa.

The production schedule is shown in Table 25-4. The schedule assumed an arbitrary project start-up of February 1, 2007. The production tonnage by quarter is shown in Figure 25-17

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Table 25-4: Ore Production Schedule

Year	Period	Ore		Cu	Au	Ag
		(000	)t	(%)	(g/t)	(g/t)
2007	Qtr 1	-		-	-	-
	Qtr 2	-		-	-	-
	Qtr 3	-		-	-	-
	Qtr 4	-		-	-	-
2008	Qtr 5	-		-	-	-
	Qtr 6	-		-	-	-
	Qtr 7	11		0.39	0.39	0.52
	Qtr 8	41		0.64	0.58	0.87
2009	Qtr 9	53		0.63	0.55	0.87
	Qtr 10	122		0.63	0.57	0.91
	Qtr 11	263		0.65	0.59	1.15
	Qtr 12	349		0.69	0.61	1.06
2010	Qtr 13	387		0.75	0.67	1.16
	Qtr 14	400		0.82	0.72	1.27
	Qtr 15	403		0.85	0.74	1.35
	Qtr 16	401		0.87	0.75	1.39
2011	Qtr 17	403		0.91	0.76	1.48
	Qtr 18	418		0.90	0.75	1.52
	Qtr 19	926		0.91	0.75	1.52
	Qtr 20	1,009		0.91	0.75	1.49
2012	Year 6	4,037		0.92	0.76	1.57
2013	Year 7	4,031		0.87	0.74	1.71
2014	Year 8	4,030		0.85	0.79	2.03
2015	Year 9	4,051		1.01	0.92	2.60
2016	Year 10	4,049		1.10	0.78	3.16
2017	Year 11	4,002		1.06	0.68	2.85
2018	Year 12	4,000		1.12	0.67	2.81
2019	Year 13	4,000		1.11	0.69	2.60
2020	Year 14	4,000		1.02	0.63	2.45
2021	Year 15	2,970		0.83	0.51	1.97
	TOTAL	44,355		0.98	0.72	2.27

Ore grades for copper, gold are shown by quarter in Figures 25-18 The estimated grade of arsenic grade is shown in Figure 25-19.

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Figure 25-17: Scheduled Tonnage from Each Block by Quarter

Figure 25-18: Scheduled Copper and Gold Grades by Quarter

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Figure 25-19: Scheduled Arsenic Grades by Quarter

25.2 Recoverability (This section was written by John Shillabeer P.Eng., Hatch)

The recovery of the mineral resources by mining has been addressed in Section 25.1. The metallurgical recovery of the mineral reserves is addressed in Section 18.

25.3 Markets and Transportation (H.M. Hamilton and Associates Inc.)

H.M. Hamilton and Associates Inc. (Hamilton) were commissioned by New Gold Inc. to complete the Transportation and Marketing components of the Hatch feasibility study on the New Afton project.

25.3.1 Markets

Production from the Project between 2009 and 2014 will be from the Hypogene zone ores and, as such, should not incur penalty charges for mercury and arsenic. In the later stages of the mine operation (2015-2021), the blending of the Hypogene and the Mesogene ores will increase arsenic and mercury to penalty levels in the concentrate produced. The marketing of these qualities would be achieved by distributing them amongst a number of selected smelters.

New Gold’s strategy will be to develop long term contracts that would include the initial high quality production along with the more complex concentrates produces in the later stages of the mine life. Smelters/Traders would be more inclined to accept the package because they would receive the higher quality in the early stages of the contract. Recently, smelters in Europe (Atlantic Copper), India, Canada (Horne) and Chile (Altonorte and Codelco) have been prepared to blend concentrates with higher levels of arsenic and mercury with concentrates containing lower levels of penalty elements. Altonorte in Chile has been the largest taker aside from Codelco's smelter.

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A forecast of the detailed concentrate assays, over the mine life, will be developed in conjunction with the completion of the next mine production schedule. At that time, smelters/traders will be able to fully respond to a request by New Gold for a Letter of Interest or a Letter of Intent for the New Afton production.

In the absence of letters of interest or letters of intent from potential smelters to further define the potential terms for the concentrates, Hamilton has generated assumptions for smelter terms with respect to treatment charges, penalties, accountability and other areas were reviewed in the light of the current market, as well as historic and future expected trends. In addition, there were discussions with smelters in order to determine current trends in terms and the long-term outlook on capacity availability and the minor element impact. With possible copper smelter outlets in Japan, China, India, Canada and Europe, the terms for a Japanese delivery have been used for comparative purposes.

Smelters in different market areas may use different formulae with respect to metal accountability and charges, which is reflected in the presented terms.

Table 25-5: Detailed Terms - Copper Concentrates

		<30%	96.5 % (minimum deduction 1.0%)
	Copper	30% - 35%	96.65 % (minimum deduction 1.1%)
		>35%	96.7 % (minimum deduction 1.1%)
	Silver		0 % of the Silver content if< 30 gms
			90%of the Silver content if> 30 gms
Payables			90 % of the Gold content if 1-3 gms
			92 % of the Gold content if 3-5 gms
	Gold		95 % of the Gold content if 5-10 gms
			96 % of the Gold content if 10-15gms
			97 % of the Gold content if 15-20gms
			97.5 % of the Gold content if >20gms
Treatment Charge			US $ 80 per dmt of concentrates CIF Free Out Japan
			US$0.08per payable pound of Copper plus when the Copper Metal
	Copper		price paid exceeds $0.90 per pound add 10% of the excess to the
Refining Charges			refining charge to a maximum of $0.10 per pound.
	Silver		US$0.35per payable ounce of Silver
	Gold		US $ 6.0 per payable ounce of Gold
	Arsenic		0.10 % free, US $ 3.00 for each 0.1 % thereafter
Penalties	Hg		20 ppm free, US $2.00 for each 100 ppm thereafter
			Moisture content: 8% free, US$1.00 for each 1% thereafter

25.3.2 Transportation

The concentrates could be delivered to Asian ports, most likely to Japan, South Korea or northern China. The Japanese destination was used as the base case for analysis. For this movement there will be three elements of cost, (i.e., a truck/rail haul from the mine to the export port of North Vancouver B.C.), storage and vessel loading at North Vancouver, and ocean transportation in bulk carrier vessels to the receiving ports. All of the receiving smelters bear the cost of unloading at the receiving ports plus any inland transportation costs to the smelter.

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25.3.2.1 Delivery Destinations

Highland Valley Copper (HVC) has two established load-out facilities in Ashcroft one on each of Canadian National Rail (CN) and Canadian Pacific Rail (CP). Only a small extension at either of these facilities would be required to handle the additional New Afton Concentrates. The bulk of the concentrates will be moved to Vancouver Wharves (VW) by CP for export. Should sales be made to eastern Canadian smelters, a rail fleet using CN would be established.

25.3.2.2 Truck/Rail Movement

The cost of the movement from the mine to VW using Arrow Transfer and CP via the HVC transfer facility at Ashcroft B.C. is C$28.75 per short wet ton (swt).

25.3.2.3 Port Handling

Storage and loading through VW has been costed at C$26.57 per wet metric ton (wmt).

25.3.2.4 Ocean Movements

Simpson, Spence and Young, a Vancouver-based shipping broker, estimate the long-term rates for shipment of concentrates from Vancouver to Japanese ports to be US$40.00/wmt.

This rate assumes vessel loadings of 10,000 wmt each.

25.3.2.5 Total Transportation Costs

Assuming an 8% moisture and an exchange rate of C$:US$ 0.88. The totals of these projected movements are, in US$ per dmt:

Truck/Rail	US$30.23
Storage and Loading	US$25.41
Ocean Freight	US$43.48
Total	US$99.12

25.4 Environmental Considerations (This section was written by Rolf Schmitt, P.Geo., Rescan Environmental Services)

25.4.1 Bond

Bonding is already in place for the current works. Going forward this will be adjusted to meet regulatory requirements.

25.4.2 Remediation and Reclamation

The initial plans for decommissioning and closure presented below have been based on the best information available at the present time, as well as feasibility level engineering design drawings. Following detailed engineering and subsequent reclamation research programs, a comprehensive and detailed reclamation and closure plan will be submitted to the BC Government for review and approval.

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The NGD 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, Ministry of Energy, Mines and Petroleum Resources 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 because of demonstrated impracticality, they have been exempted by the inspector

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

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 immediately. 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 immediately.

The side slopes of the tailings storage facility will not be reclaimed until closure. However, it 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.

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. Sparse vegetation may indicate compaction.

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 in a book which will be kept on-site.

The reclamation of the tailings storage dam 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 immediately. Records will be kept to assess the changes in percent bare ground or surface cover as well as plant health.

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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 BC 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.

Materials will be moved, stockpiled, and stored on the site during the construction and operation phases. During construction, the pipeline corridor and access roads will also be subject to light fertilization and low application rates of revegetation mix to stabilize exposed soils. In Year 1 – 5 the Pothhook Pit will be filled with tailings and reclaimed. The Pothook Pit earth dam will be dismantled. Upon closure, the plant facilities will be removed and the tailings storage facility closed. The closure of the tailings will allow for final reclamation of the dam slopes and surface. The visual berm will be dismantled and the plant site reclaimed with the material for the berm. The detailed reclamation activities over the duration of the mine operation are presented in the Mine Permit Application, along with a detailed costing of activities.

Table 25-6 provides a summary of the infrastructures which require revegetation and reclamation.

This table includes the infrastructures which require short term and long term treatment. 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 tailings storage facility dam which comprises 37 ha of surface and access road slopes which comprises 4.6 ha, will be reclaimed. Therefore, a total surface area of 154.1 ha will have received final reclamation with the remaining infrastructures treated to prevent wind and water erosion during the construction and operation of the mine.

Table 25-6: Summary of Infrastructures Requiring Reclamation

Infrastructure	Slope	Type Seed	Fertilizer	Hydroseed	Footprint	Volume (m3)	Surface
					(ha)		Area (ha)
Topsoil Stockpile	Sloped	X				162,500	2.0
Overburden Stockpile	Sloped	X	X			325,000	3.2
Pothook Pit Dam	Sloped	X	X			37,000	1.3
Tailings Dam	Sloped			X			3.7
Visual Berm	Sloped	X	X			6,000	0.8
North Stockpile	Sloped	X	X			53,200	1
Access Road	Sloped	X	X				4.6
Pothook Pit	Flat	X	X		10.2
Pothook Pit Dam	Flat	X	X		1.2
Tailings Surface	Flat	X	X		83.1
Plant Site	Flat	X	X		13
Plant Site Subsidence Zone	Flat	X	X		5
Total					112.5		49.9

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25.5 Taxes (This section was written by NGD and reviewd by PricewaterhouseCoopers LLP)

Independent accountants, PriceWaterhouseCoopers LLP have confirmed the validity of the tax treatment in the economic model.

25.5.1 Royalties

NGD is currently liable to pay certain royalties upon achieving commercial production. NGD intends to purchase the royalties and the sums required to do this in accordance with the agreements are included in the project capital. Accordingly the project assumes that upon reaching commercial production no royalties would be payable.

25.5.2 Income Taxes

Income earned by the Afton mine is subject to federal, provincial and BC mining tax.

Capital expenditures have been reviewed and allocated amongst the following tax deductible categories with the following tax deductible rules:

Tax deductible capital expenditures are segregated for tax purposes between depreciable and depletable assets.

Treatment of depreciable and depletable expenditures prior to commercial production being attained.

Depletable assets are included in Canadian exploration expense (“CEE”) and are generally deductible up to 100% of unclaimed balances but cannot be deducted to create or increase non-capital loss. Depreciable assets are included in the tax deductible Class 41(a) and deductible up to the income from the mine for that year.

Treatment of depreciable expenditures post commercial production related to the expansion of the mine.

Depreciable expenditures incurred after commercial production has been attained which relate to an increase in the initial mill and mine design in excess of 25% also qualify for accelerated depreciation. The Projects initial design capacity is nominally 1.6 million tonnes per year and is then increased to 4 million tonnes per year during the first two years of operation. The depreciable expenditures related to the expansion are treated in the same manner as Class 41(a) as noted above.

Treatment of depreciable and depletable expenditures post commercial production excluding expenditures qualifying as mine expansion expenditures.

Depletable assets are included in Canadian development expense (“CDE”) and are deductible up to 30% of the unclaimed balance on a declining balance basis. Depreciable assets are included in tax deductible Class 41(b), subject to expenditures in excess of 5% of gross revenue, which are treated as Class 41(a.1) and deductible in the same manner as Class 41(a) noted above. Class 41(b) expenditures are deductible as a capital cost allowance up to 25% of the undepreciated capital cost (“UCC”), subject to a half-year rule for asset acquisitions during the taxation year.

The ordering of deductibility has been prioritized as CDE, Class 41(b), Class 41(a)/41(a.1) and CEE.

Non-capital losses can be carried forward for 20 years.

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25.5.3 Large Corporation Capital Tax

The Large Corporations Tax has been repealed and is no longer applicable subsequent to 2007 and has not been included in the calculations.

25.5.4 Federal and BC Income Tax, and BC Mining Tax Rates

The current legislated federal income tax rates applicable for the project life are:

		Federal		Provincial		Total
2009		20	%	12	%	32	%
2010	and hereafter	19	%	12	%	31	%

25.5.5 BC Mining Taxes

BC mining taxes are calculated at the greater amount of the 13% mining tax and the minimum 2% net proceeds tax.

BC mining taxes allow the deduction of depreciable and depletable capital costs against the net revenue from the operation for mining tax purposes. The depreciable and depletable costs are accumulated in a cumulative expenditure account (“CEA”) and any unclaimed amount can be carried forward indefinitely. Prior to commercial production, expenditures in the CEA account are subject to a one-third bump.

The minimum net proceeds tax is based upon 2% of the mine’s net proceeds (revenue less operating costs).

25.5.6 Flow-Through Funding Eligibility

The Company applied for and received a tax ruling from the Canadian Revenue Agency confirming that certain expenditures applicable to the further exploration and qualifying development expenditures of the new underground mine, subsequent to making a decision to bring the mine into production in reasonable quantities, would qualify as CEE. CEE expenditures qualify for renouncement to purchasers of the Company’s flow-through share offerings.

The above referenced income tax calculation has not considered flow-through equity raisings as a source of funding for the mine. If they were, the tax-deductible base available from the CEE tax pool would be lower by the amount renounced to the purchasers of the shares.

25.6 Capital Cost (This section was written by John Shillabeer, P.Eng., Hatch)

25.6.1 Introduction

This section was written primarily by Hatch but with significant contributions from AMC, other consultants and the Company (as described both below and in Table 4-1). It provides a summary of the project capital costs and explains the basis of the estimates, conditioning assumptions, inclusions and exclusions. Working capital, sustaining capital and closure costs are also discussed.

The project capital cost estimates for the mine, process plant and infrastructure were prepared in accordance with standard industry practices and the estimate is classified as a Class 3 estimate as defined by AACE (American Association of Cost Engineers) –5+15% accuracy. The capital cost estimate used a base date of January 1, 2007 and there is no allowance for escalation beyond this date.

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AMC estimated all mine capital costs, except those estimated by Hatch. Hatch estimated the following mine underground capital costs: water supply and distribution, main dewatering system, electrical distribution and underground communications and control. Vector estimated the capital cost of the earthworks components of tailings disposal. NGD estimated the Owner’s costs. Rescan estimated re vegetation costs and environmental monitoring costs. MEG estimated the costs of pit debris stabilization. Urban Systems estimated the cost of rehabilitating the pump station or pipeline. Hatch estimated all other costs. Hatch did not verify AMC’s, Rescan’s, MEG’s, Urban System’s nor Vector’s estimates, other than to confirm their overall fit and conformance to scope.

Representatives from Hatch, AMC and Vector visited the site and Kamloops to view local conditions, obtain prices and assess local capabilities. In addition, Hatch referenced its recent experience managing the construction of another similar processing facility in South Central BC.

25.6.2 Estimate Summary

This section contains summaries of the estimated capital cost of initially constructing and then expanding the project. Note that the estimates are subject to the exclusions, inclusions and assumptions which follow.

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Table 25-7: Capital Cost Estimate Summary by Area (C$ ‘000)

		Responsible	Project Phase
WBS	Description	Party	Initial	Expansion	Total
A0	Site Devel./Roads	HATCH1	10,361	0	10,361
F0	Process	HATCH	56,300	20,749	77,049
G0	Conc. Transfer	HATCH	148	0	148
H0	Elec. Power	HATCH	5,550	0	5,550
J0	Tailings and Waste Disp.	HATCH2	5,721	9,385	15,106
K0	Surface Serv. Facilities	HATCH	10,594	3,320	13,914
M0	Mining	AMC3	125,902	78,673	204,574
		HATCH	15,536	3,668	19,204
	Capitalized Operating Cost	AMC	2,522	0	2,522
P0		HATCH	811	0	811
Z1	Engineering	AMC	4,891	2,430	7,321
		HATCH	12,041	2,949	14,990
Z2	Procurement	AMC	1,223	608	1,830
Z3	Const. Mgmt.	AMC	6,113	3,038	9,151
		HATCH	9,918	2,676	12,594
Z4	Const. Indirects	AMC	4,700	0	4,700
		HATCH	6,905	2,826	9,730
Z5	First Fills	HATCH	291	337	628
Z6	Spares	AMC	0	600	600
		HATCH	1,251	721	1,972
Z7	Duties	HATCH	0	0	0
Z8	Freight	HATCH	2,694	1,137	3,831
Z9	Commissioning Costs	HATCH	876	688	1,564
Z10	Owner's Costs Requiring Contingency	HATCH	6,261	1,071	7,332
	Sub Total		290,607	134,875	425,482
	Contingency	HATCH	36,532	17,729	54,261
	Total		327,139	152,604	479,743
	Defined Owner's Costs (no contingency)	NGD	805	14,000	14,805
	Project Total Capital		327,944	166,604	494,548

1	Includes estimate by Urban Systems of the cost to refurbish the water supply pipeline and pumps.
2	Includes estimate by Vector of the cost of constructing the Pot Hook dam and the tailings storage facility.
3	Includes estimate by MEG of the cost of stabilizing the debris at the bottom of the Afton pit.

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Table 25-8: Capital Cost Estimate Summary by Commodity (C$ ‘000)

		Responsible	Project Phase
Code	Description	Party	Initial	Expansion	Total
A	Earthworks/Site Prep.	HATCH	10,474	4,381	14,855
B	Detailed 'works (Mining)	AMC	125,902	78,673	204,574
B	Detailed Earthworks	HATCH	2,097	196	2,294
C	Concrete	HATCH	9,418	2,196	11,614
D	Structural	HATCH	2,138	796	2,934
E	Cladding and Roofing	HATCH	0	60	60
F	Architectural	HATCH	952	591	1,543
G	Manufactured Buildings	HATCH	12,337	2,006	14,343
H	HVAC	HATCH	624	75	699
M	Process Mechanical Equip.	HATCH	31,065	14,994	46,059
N	Mobile Equipment	HATCH	2,060	0	2,060
P	Piping	HATCH	3,202	2,229	5,431
Q	Platework	HATCH	1,724	520	2,244
R	Pipelines	HATCH	6,764	4,214	10,978
U	Electrical Power and Control	HATCH	18,022	4,687	22,709
V	Instrumentation Control	HATCH	2,003	177	2,180
W	Communications	HATCH	1,329	0	1,329
	Capitalised operating utilities	HATCH	811	0	811
Sub total: Direct Costs			230,922	115,795	346,718
Z	Indirect costs	HATCH&AMC	59,685	19,080	78,765
Contingency			36,532	17,729	54,261
	Defined Owner's Costs (No Contingency)	NGD	805	14,000	14,805
Project Total Capital			327,945	166,604	494,549

* Some estimates include estimates by others as explained in the footnote to Table 25-7.

NGD has advised Hatch that the cost of acquiring the surface rights from Teck (C$17,030,000 including interest) is not part of project capital and instructed Hatch that it should be excluded from the above tables and the economic analyses.

25.6.3 Basis of Estimate

25.6.3.1 Base date

The base date of the estimate is January 1, 2007

25.6.3.2 References

The references used for estimating the capital costs include: drawings, equipment lists, project schedule quotations for all equipment shown on flow sheets, electrical equipment shown on single line diagrams (preliminary quotations), mobile mining equipment, in-house data for similar recent projects, bulk material quantity take-offs, site productivity review and assessment, labour cost surveys and assessment, sub consultants’ reports, and local construction contractors’ budget quotes.

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25.6.3.3 Exchange Rate

All costs shown in the estimate are in Canadian dollars (CAD). The following conversion rates were used

$1 CAD =US$0.85

$1AUD = C$0.80

25.6.3.4 Mining (This section was written by Mike Thomas, MAusIMM (CP), AMC)

AMC estimated the mine capital cost as an Owner-developed mine; that is as if NGD employed all the workers and staff, purchased and maintained the equipment. NGD intend that the mine will be developed and initially operated by a contractor. The contractor will work within the budgets derived form this estimate.

AMC based mine labour costs on shift employees working three, eight hour shifts on a continuous system with each employee’s payments based on 2190 paid hours per year. The technical and managerial staff will work a conventional five day week. Wage rates and salaries were developed from the western region component of a survey of Canadian salaries and wages, selecting rates based on a point 75% of the way between the survey minimum and maximum rates. AMC also added a shift allowance of $0.82. Salaries were based on a published 2006 survey of Canadian mining industry salaries prepared by Price Waterhouse Coopers.

AMC developed average hourly rates for typical excavation and construction crews and used these average hourly crew rates to estimate the costs of development and underground construction. Specialist contract labour required for installing civil, mechanical and electrical components of the conveyor and crushing system were costed at the rate of $64 per hour. This rate is consistent with the construction labour rates used by Hatch for surface works and is higher than the rates for underground mining because it includes allowances for supervision, small tools, temporary facilities, contractor’s overhead and profit.

AMC based their capital cost estimates on the assumption that the mine mobile equipment will be used for the initial mine development, ongoing development and block cave production. During initial development a fleet of large drilling and mucking equipment will be supported by 50t trucks hauling waste to surface. As the extraction and undercut levels are developed, smaller equipment will be introduced and some of the larger equipment will be withdrawn. The capital cost estimate is based on all new equipment purchased specifically for the project and no credit is taken for the resale of equipment that becomes redundant, such as the 50t trucks. This is conservative. NGD also has the option of leasing mine equipment. AMC developed hourly operating costs for mine mobile equipment, applied them with appropriate productivities and physical quantities to obtain a unit cost for each activity (e.g. drilling, mucking, truck haulage, etc). They then applied the unit costs to the quantity take-offs from the development schedule to obtain the estimated cost of activities during each specific period.

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AMC derived development costs per meter from first principals for each type of development. The costs comprise blast hole drilling, blasting, loading to stockpile bays, ground support and mine services but exclude truck loading and haulage, and operating and maintenance labour costs. Truck loading and haulage was estimated separately depending on distance. Table 25-9 summarises the average cost per meter according to the WBS structure. These costs include truck loading from stockpile bays, truck haulage and labour. However equipment ownership costs, i.e., costs associated with the capital purchase (or lease) of the equipment, are excluded.

Table 25-9: Lateral Development Costs by WBS

WBS	Description	Average cost (C$/m)
M1	Mine accesses	3,770
M2	Undercut level development	2,547
M3	Extraction level development	4,081
M4	Ore transfer level	4,055
M6	Conveyor decline	4,729

AMC estimated the costs of main ventilation raises using contract raise bore rates obtained from Cementation. The rates include rig set up, pilot hole drilling, reaming and rig removal. AMC developed from first principals the costs of mining short raises for ore passes and ventilation at the ends of the extraction level crosscuts. Rates for vertical development are summarised in Table 25-10.

Table 25-10: Vertical Development Rates (excluding muck removal)

Type	Dia	Cost
	(m)	(C$/m)
Raise bore	3.5	4,293
Raise bore	2.4	3,818
Longhole raise	3.5	533
Longhole raise	2.4	369

AMC estimated the cost of establishing the undercut and drawbells from first principals, based on indicative drill patterns and estimates of the costs of drilling, consumables, explosives, maintenance and equipment productivity rates. A summary of cost estimates is shown in Table 25-11.

Table 25-11: Drawbell Drilling and Blasting Costs (excluding labour)

Type	Drilling cost	Basting cost
	(C$/m drilled )	(C$/m drilled)
Undercuts drill and blast	9.83	14.83
Drawbells drill and blast	9.82	10.67
Slot raises (V30)	220.25	-

The cost of supplying and installing steel sets and additional support work in constructing a draw point is estimated at $24,224 per drawpoint. This cost excludes labour.

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AMC based their estimate of the cost of components of the crusher station and the conveyor system on material take offs from the feasibility study design for civil, structural, mechanical and electrical components. The costs of the components were obtained from quotations in Canadian dollars. In certain cases Australian prices were used and converted to Canadian dollars at the appropriate exchange rate. Installation costs were estimated for the crusher station, based on a percentage of the total cost of components with the percentage derived from AMC’s experience on other similar installations. The installation cost of the conveyor system was estimated by developing an estimate of the labour required for installation and by applying factors to the cost components for the hire of equipment and specialist contractors.

The cost of components of the main surface fans and the heating plant were estimated by AMC based on quotations from Canadian suppliers. Installation costs were based on typical costs for horizontally mounted fans of the type proposed. The cost of secondary fans and other ventilation control devices was based on costs provided by Canadian suppliers.

Hatch estimated the cost of the main pump station. AMC estimated the costs of the secondary pump station, pumps, sumps, pipe work, other materials required for development and miscellaneous face dewatering based on quotations obtained from US suppliers of this equipment. Prices were converted to Canadian dollars at the appropriate exchange rate.

AMC included estimates for the cost of recruiting, training, personal and other safety equipment, computers, software and miscellaneous technical equipment, office consumables, external technical support, geotechnical monitoring, control systems and road maintenance. AMC included an allowance for the cost of geological and geotechnical drilling required to provide information for the final detailed mine design. They included a total of 8,000 m of drilling to complete the design of B1 and B2 and 4,000 m for B3. A further allowance of 4,000 m was included to complete the detailed design of the conveyor decline, crusher chamber and ventilation shafts. AMC applied a unit cost of $120/m inclusive of contract drilling, logging and assay.

25.6.3.5 Surface Infrastructure and Process Plant (Hatch)

Earthwork quantities for the process plant and surface infrastructure were developed from topographic plans, drawings and sketches. This includes the service roads within the plant site, the access road and site preparation. Unit cost rates developed from local contractors were applied. Site fill and engineered fill were based on locally available materials

Concrete quantities were developed from mechanical general arrangements, concrete design engineering and layouts drawings and sketches. Concrete unit rates were based on local cost information and were all inclusive of formworks, reinforcing, concrete and other materials, equipment, installation labour, and contractor’s overhead and profits. Excavation for foundations was excluded from the concrete unit prices. Excavation was included under earthwork.

Structural steel quantities were developed from general arrangements, preliminary structural drawings, design calculations and sketches. Structural unit rates were based on local cost information and were inclusive of material, installation labour, contractor’s indirects and profit.

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Structural platework requirements for bins, tanks, hoppers, chutes, pump boxes and launders were developed from sketches and all inclusive unit rates were applied.

The capital cost for the main process, shop and warehouse buildings were based on budget quotations for pre-engineered type buildings meeting short form specifications and local building codes. In each case, suppliers quoted on a supply and erect basis. Hatch added the estimated cost of the foundations, additional structural steel, electrical and mechanical services, interior fittings and furniture.

The office and dry complex is modular and the cost was based on a quotation selected from among several when Hatch sent suppliers the original conventional building design and short form specification. Candidate vendors of the modular office dry complex were permitted to propose their own solutions, provided that they met the original intent. Quotations were provided on a supply and erect basis. Hatch added the costs of preparing the site, providing services and furniture.

The cold storage building is a “Sprung” type structure and the estimated cost was based on a supply and erect quotation. Hatch added the cost of the local site preparation and foundation.

Equipment requirements were based on the equipment list and process flow diagrams. Vendor budget costs were obtained based on preliminary specifications and data sheets for all items of mechanical equipment shown on the process flow diagrams. Full specifications were written for the SAG and ball mill.

Costs for installation of equipment were based on unit work hour requirements to install the equipment adjusted for local conditions and unit rates developed for process equipment installation

Quantities for electrical work were developed from the overall system single-line diagram, general arrangement drawings, equipment list and process flow diagrams. Vendor budget quotes were obtained for major equipment. Supply and installation of material were based on labour estimates and in-house historical data for similar projects.

Power requirements for equipment, as indicated on the equipment list, were used to run a load analysis. From the load analysis, main power distribution components such as MV and LV switchgear, main cables and transformers were sized. The estimate was based on tying in to the existing BC Hydro 138 KV power line.

The control system is a programmable logic controller (PLC) with operator interface stations. Budget quotations were obtained for PLC and HMI based on number of I/O’s. Supply and installation of material was based on labour estimate, budget quoted prices and in-house costs. Costs for wiring local instruments were factored based on in-house data.

Process and service piping quantities were developed based on process flow diagrams, the general plant layout and factors developed from a current similar project now being built. Costs of pipe, fittings and valves were based on recent in-house data. All yard piping, tailings disposal and water reclaim piping was based on material take off quantities and layout drawings.

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Construction labour estimates were based on non-union labour rates and productivities. Hatch developed a current schedule of pay rates and developed the fully loaded rates shown in Table 25-12, considering local conditions of employment and current experience on another similar project nearby. Hatch estimators also obtained advice from Kamloops area contractors.

The fully loaded labour rates shown in Table 25-12 include all benefits, contractor site supervision and administration, mobilization, demobilization, small tools and consumables, contractor temporary site facilities and services, and contractor overhead and profit.

Table 25-12: Construction Labour Rates

Activity	C$/h
Earthwork	56
Concrete	61
Steel plate work	62
Steel structural	63
Architectural and Building	61
Mechanical	63
Piping	62
Electrical	62
Instrumentation	65

Construction indirect estimates recognised the cost of assuming that 50% of the work force will come from elsewhere. This cost was not included in the labour rates shown above but was separately estimated. Also, these rates excluded construction equipment. This also was separately estimated.

Installation work hours were based on in-house historical data, adjusted for various crafts to reflect local site conditions.

25.6.3.6 Tailings Management (This section was written by Vector)

Vector Engineering Inc designed the tailings disposal systems. Hatch designed the pumping and piping to suit. Vector provided design drawings, calculated quantities (material take offs) estimated the availability of local construction materials (i.e., soils ) and estimated the costs of construction using in house (US) civil construction costs. Hatch compiled Vector’s cost estimates, converted them to Canadian dollars and calibrated them with local Canadian rates.

25.6.3.7 Pit Debris Stabilization (This section was written by MEG)

MEG devised a method for stabilizing the now-saturated debris located in the bottom of the Afton open pit. Their provisional approach is described in Section 25.1.3. MEG provided preliminary estimates of the cost of their road building, well drilling and well pumping program which Hatch augmented with local civil contractor labour and pipe installation costs. The MEG program costs are considered preliminary. The program will be defined in greater detail after the pit has been dewatered and the debris becomes accessible for sampling and measurement.

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25.6.3.8 Assumptions

The following assumptions are part of the basis of the estimate:

• The project will be built in accordance with the scope, schedule, drawings, specifications quantity estimates and execution plan contained in this report.

• 50% of surface construction labour is imported and 50% is available locally. Only the non local labour is paid a living out allowance of $120 per day

• The project is based on the procurement of all new equipment.

• Riaseboring is done by a specialist contractor who provides his own equipment. Otherwise, all mine development is completed with equipment that is purchased for the project.

• The mine development cost estimate and schedule (based on Owner direct hire) is similar to that achieved by the selected contractor working under an open book risk sharing contract.

• The project is PST exempt

• One apprentice will be employed for every five tradespersons underground

25.6.3.9 Exclusions and Inclusions

Following are part of the basis of estimate:

25.6.3.9.1 Exclusions

The following costs and cost generating activities are excluded from this estimate:

• Surface adits acquisition cost.

• Operating costs which are not capitalised

• Financing costs

• Interest during construction

• Performance bonds

• Special allowances (e.g., for overbuy and scope creep)

• The cost of settling claims

• Used equipment acquisition, inspection and refurbishment costs

• Facilities not identified in this report

• Any environmental, archaeological and ecological considerations beyond those expressly included in the present design.

• Any costs incurred in connection with BC Hydro’s generating capacity or BCTC’s transmission system due to the addition of the New Afton power demand profile.

• Any and all costs to accelerate the work beyond the present schedule which is based on six days a week, ten hours a day for surface construction and three eight hours shifts per day continuously underground.

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• Sunk costs – all costs incurred up to the time the feasibility study is submitted.

• Currency exchange fluctuations (exchange rates used are reference in Section 25.6.3.3)

• Price escalation beyond January 1, 2007

• Costs incurred to dispose of existing hazardous soils or other substances.

• Import duties and GST

• Changes to legislation, regulations and policy.

• Costs that may be incurred as a result of the failure of New Gold to follow the project execution plan described in this report

• Any additional work that is required as a result of subsurface conditions in and around the project site that were not known as of the base date of the estimate.

In addition no allowance has been made in the capital estimate for any of the following risk factors:

(a)	project risk factors that would be expected to potentially impact any project such as this Project (e.g. adverse weather conditions, acts of God and other force majeure events, delays due to unforeseen factors such as late delivery or unavailability of equipment or materials, or unavailability of labour resources, poor performance by EPCM contractors or construction contractors, disputes with local residents, etc)
(b)	the project-specific risk factors identified in risk management workshops and documented in the feasibility report. Or
(c)	political legal or regulatory risk factors (e.g. changes to laws, changes to taxation or royalty regimes, or non issuance, cancellation or revocation of permits or licences required to develop and operate the project).

Any costs that may be incurred as the result of the occurrence of any of these risk factors have not been included in the capital cost estimate.

25.6.3.9.2 Inclusions

The following costs and cost generating activities are included in the estimate:

• Allowances for the costs of items that are too small to individually estimate

• A reserve for contingencies

• Permits

• Detailed engineering design and procurement

• Technical assistance during commissioning and construction

• Temporary facilities

• Equipment use din construction

• Project insurance

• Property taxes as specified by the owner

• Initial stores and supplies

• Freight

• Security

• Community support and liaison

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• Training and operating manuals

• The cost of purchasing the net smelter return royalties

• Mobilization and demobilization

• Construction management

• Other Owner’s costs attributed to the project

25.6.4 Working Capital

Working capital was estimated in the economic model. Table 25-13 provides a summary of working capital requirements during the project. Additional working capital will be required in 2012. In subsequent years some of the working capital can be withdrawn.

Table 25-13: Working capital invested during the project (US$ 000's)

	2009	2010	2011
Total Working Capital Balance	9,599	21,831	27,050
Working Capital Invested	9,599	12,232	5,227

25.6.5 Sustaining Capital

After the mine expansion is completed, capital will be required to sustain the business by developing Block 3, extending certain underground services and replacing mobile equipment. Capital will not be required for tailings disposal, nor process equipment replacements. Table 25-14 provides a summary of sustaining capital requirements. All of the sustaining capital was estimated by AMC.

Table 25-14: Sustaining capital (C$ ‘000)

WBS	Description	2011	2012	2013	2014	2015	2016	2017	2018	2019	2020	Total
M1	Mine Access	-	-	-	2,290	226	-	-	-	-	-	2,516
M2	Undercut	2,194	4,791	4,299	2,636	6,037	4,910	2,395				27,263
M3	Extraction Level	3,392	5,629	6,027	4,963	5,971	10,676	3,718				40,377
M4	Ore Transfer	176	90	277	201	43			-	-	-	788
M8	Ventilation	154	279	85	2,274	408	116	-	-	-	-	3,316
M10	UG Pumping & Drainage	-	-	-	-	246	-	-	-	-	-	246
M14	Mining Mobile Equipment	-	-	400	2,082	475	4,897	-	-	1,302	-	9,156
M17	Geology & Diamond Drilling	-	-	-	480	-	-	-	-	-	-	480
	Total	5,917	10,790	11,088	14,926	13,406	20,600	6,113	-	1,302	-	84,142

25.6.6 Closure Cost

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 re-grading the site. The resale value of equipment and buildings was ignored. All underground equipment was left in place.

Rescan estimated the cost of re distributing stockpiled till and top soil and of revegetation. They also assessed the cost of addressing remaining legacy liabilities from the previous operator and estimated the cost of post closure monitoring. All of the closure cost estimates were prepared in accordance with MEMPR guidelines and used the MEMPR spreadsheet. Table 25-15 contains the closure cost estimate.

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Table 25-15: Closure Cost Estimate

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
Sub Total			$	7,056
Reclamation (Rescan)
Tailings Dam				154
Road				2
Pothook Pit				4
Pothook pit earth dam footprint				0
Tailings surface				3,432
Plant site				299
Sub Total			$	3,891
Round off				53
Total Cost			$	11,000

 

The closure cost has been scheduled to be spent in equal amounts in 2021 and 2022.

25.6.7 Contingency

25.6.7.1 Introduction

Contingency is a cost element to cover unknown items that are expected to occur within the defined scope of the project but which cannot be properly defined at the feasibility stage of the project. It should be assumed that the contingency will be spent.

The contingency allowance specifically excludes costs arising from scope changes, project risk factors and all other times that are excluded from the capital cost estimate (see Section 25.6.3.9.1) .

While it is expected that contingency funds would be spent, they are intended to cover only those factors that influence the cost for the currently defined scope, including:

• Inaccuracies and minor omissions to project scope definition;

• Any specific unavoidable cost increases

• Routine minor changes of a non-fundamental nature that occur during project execution.

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25.6.7.2 Estimate Accuracy Contingency

Each line item of the capital estimate was assigned a unique code (using a common coding system) that reflects the following:

• The level of confidence in the estimated cost

• A distribution of possible variation in the estimate

• The possibility that the estimate may be affected by a future scope change

• The probability of and degree to which the estimate may be erroneous due to the inherent complexity of that being estimated.

A Monte Carlo statistical simulation was then conducted on the capital estimate using @Risk software. Inputs to the simulation included the individual line items and associated codes. Output from the simulation was a probability curve illustrating a range of probable capital costs for the project against the probability (confidence) of achieving the costs. A confidence level was then selected depending on the Owner’s tolerance of risk in order to derive the contingency allowance. The simulation analyzed the initial project and then the expansion to define separate probability curves for each phase of the project. As the analysis was conducted on the project’s phases as a whole, it was not relevant to generate contingency amounts against each WBS element.

Based on the Monte Carlo simulation analysis described above, there is a 80% confidence that the cost of Phase 1 would not exceed the base estimate plus a contingency allowance of 12.6% and that the cost of the Phase 2 would not exceed the base estimate plus a contingency allowance of 13.1% . The project team recommends that if there is to be an 80% probability that the budget will not be exceeded, the Phase 1 budget should be the base estimate plus 12.6% and the Phase 2 budget should be the base estimate plus 13.1% .

25.7 Operating Cost (This section was written by John Shillabeer, P.Eng., Hatch)

This section was compiled by John Shillabeer (Hatch Ltd.) and includes mine operating cost estimates provided by Mike Thomas (AMC); processing, surface general and administrative estimates prepared by Ken Major (Hatch Ltd.).

25.7.1 Summary of Mine Life Operating Costs

The cost estimates reported in this section are not escalated. The effective date of the estimate is January 1, 2007.

Figure 25-20 illustrates the variations during the project of the overall unit operating cost per tonne milled and per pound copper. These costs include off-site transportation and processing of concentrate. Costs are highest in the initial years when operating at 1.6 Mt/y because all ore must be hauled to the surface in trucks, crushed in a temporary crusher operated by a contractor and re handled using a front end loader before it is introduced into the SAG mill. Economies of scale become apparent when output is increased to 4 Mt/y in 2011 and the underground crusher and conveyor system replaces the trucks. Tailings disposal costs are moderate in the first two years, compared to the expanded operations phase, because the whole tailings are simply placed in Pothook pit requiring minimal operating labour.

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Figure 25-20: Life of Mine Unit Operating Cost

Somewhat elevated ore grades in the first two years help to moderate the higher operating costs until production is expanded to 4 Mt/y.

Table 25-16 summarizes average unit operating costs over the life of the Project, (may not add to totals shown due to rounding). Costs shown in the table are on-site operating costs per tonne milled (they exclude concentrate transport and processing). Note that the table of mining costs, Table 25-18 quotes costs per tonne mined.

Table 25-16: New Afton Total Operating Unit Costs

C$ Per tonne milled	Initial	Expanded	Life of Mine
Mining	11.80	4.61	5.14
Processing	7.98	4.02	4.33
G& A	2.03	0.99	1.08
Utilities	3.10	2.15	2.23
Total	24.90	11.76	12.76
US$ Per pound of payable copper
Mining	0.64	0.21	0.23
Processing	0.43	0.18	0.20
G & A	0.11	0.04	0.05
Utilities	0.17	0.10	0.10
Total	1.34	0.53	0.58

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25.7.2 Basis of Estimates

25.7.2.1 Base Date

The base date of the estimate is January 1, 2007

25.7.2.2 Methodology

Estimators prepared operating cost estimates by synthesis of operating and maintenance labour productivities, supplies consumption and energy consumption, based on experience and by benchmarking against other operations, where appropriate. Operating costs for all defined activities in a period were assembled in a master template and reviewed prior to becoming inputs to the economic model.

Salaries ,wages and benefits were determined from references⁵⁶. A standard set of labour and salary rates were derived from these reports. All labour and salary rates used were fully loaded for burden (i.e. there was not a separate cost centre in which all the fringe benefits were pooled).

AMC estimated the mine operating cost as an Owner-operated mine; that is as if NGD employed all the workers and staff and operated and maintained the equipment. NGD intend to that the mine will be initially operated by a contractor and that the contractor will be expected to work within budgets derived from this estimate.

The mining workforce was estimated by determining the labour required to carryout the various activities (mucking, truck haulage, crushing/conveying, etc) required at different stages throughout the mine life. To do this AMC used its knowledge of productivities for similar activities at other mining operations. The same approach was used to estimate maintenance labour, supervision, management and technical staff. Labour costs for each activity were then estimated by applying the wages and salaries, including benefits, estimated in the manner described in Section 25.6.3.4.

Mobile equipment engine operating hours (usage) were estimated for the key items of mobile equipment. Hourly operating costs, developed from cost and performance data provided by equipment suppliers and from AMC’s benchmarking database, were then multiplied by the usage to estimate the equipment operating costs for each activity. Allowances were made in the estimate for accidental equipment damage.

The quantities and cost of parts and materials, including replacement conveyor belt, required to operate the conveyor were estimated and averaged over the planned operating life of the conveyor system. The annual cost of parts and materials to maintain and operate pumps, crusher, ventilation equipment and the electrical and communication systems were estimated as a percentage to the purchase and installation cost of the equipment. The percentages applied have been based on AMC’s experience of typical operating and maintenance costs for each equipment type.

AMC provided advice to Hatch on electric motor sizes and the expected equipment utilisation for the mine, including the ventilation system. AMC also provided an estimate to Hatch of natural gas consumption for ventilation air heating costs.

⁵ 2006 Mining and Processing Plant Salary and wages Survey: Western Mine Engineering

⁶ 2006 Mining Industry Salary Surveys Site Report-Canada, July 2006: Price Waterhouse Coopers

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Estimates were included of the costs of recruitment, training, personal and other safety equipment, computers, software and miscellaneous technical equipment, office consumables, external technical support, geotechnical monitoring and control systems, Allowances have also been made in the estimate to maintain the underground workings and repair damaged draw points and ore passes.

General and administration costs were estimated from the organisation chart (for labour and salaries) and by experience with other operations for non labour costs.

The cost of electrical energy used per period was estimated from the master load list by applying suitable values for the utilization, diversity of demand and efficiency factors for each piece of equipment. The overall operating power cost was calculated (by Hatch) with separate sub totals for the mine, processing plant and general and administrative areas.

25.7.2.3 Key Commodity Costs

The following key commodity costs were used in the estimates:

Table 25-17: Commodity and utility costs used in operating cost estimates

Electrical energy7:
Demand charge	C$4.866 per KVA per month
Energy charge	C$0.02852 per KWh
Natural Gas	C$10.28/GJ plus $124.95 per month
Diesel Fuel	C$0.80/litre
Fresh water	Nil, but operating costs include the cost of operating the pumps and pipeline from Kamloops Lake

25.7.3 Assumptions

• The mine will be built and operated in line with the relevant chapters of the feasibility study.

• The mine will be operated by a contractor until after the expansion project is completed and the ramp up in production begins.

• The mill will be operated by NGD employees throughout.

• The production schedule will be as shown in the feasibility study.

• Ground and other natural conditions will be as projected and in line with tests conducted to date.

In preparing the operating cost estimate it has been assumed that none of the events excluded from the capital cost estimate occur. The occurrence of any of these events (e.g. scope changes, project risk factors, changes in laws) might have a material impact on operating costs.

⁷ BC Hydro Tariff Schedule 1823, effective July 2006

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25.7.4 Exclusions

• All costs that are classified as capital

• Contingency

• Escalation

• Exploration

• Depreciation and amortization

• Interest

• Income taxes

• Royalty payments

• Off site transportation of concentrates (classified as part of the net smelter return and deducted from gross revenue)

• GST

• Provincial sales tax (this is an approximation)

25.7.5 Inclusions

• Head office costs attributable to the operation

• Replacements of equipment and machinery that cost less than $500,000.

• Allowances as appropriate for dilution, losses, wastage.

• Insurance

• Property taxes

• Freight

• Training

25.7.6 Mining (This section was written by AMC)

Operating cost estimates for mining were provided by AMC Consultants. The total mine operating cost estimate for the project shown in Table 25j-18 and Figure 25-21 was based on the following assumptions:

• Mobile equipment used to operate the mine will be purchased new and the cost to acquire it is included in the capital estimate.

• All costs associated with forming the undercut, including longhole drilling and blasting have been capitalized and are therefore excluded form the operating cost estimate. All extraction level development including drawpoints and drawbells have been handled in a similar manner.

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Table 25-18: Average Mine Life Unit Operating Costs Excluding Electric Power and Heating (C$/tonne mined) (AMC)

	Initial	Expanded	Life of Mine
Extraction level operations	2.78	1.55	1.63
Transfer level operations	0.07	0.54	0.50
Haulage level operations /truck haulage	6.44	0.58	0.98
Underground crushing	0.00	0.22	0.20
Conveying to surface	0.00	0.48	0.45
Maintaining underground openings	0.17	0.05	0.06
Ventilation (primary and secondary)	0.15	0.07	0.07
Dewatering	0.21	0.09	0.10
Supplies transport and other support	0.64	0.23	0.26
Mine management and technical support	2.22	0.81	0.90
Total	12.68	4.61	5.14

Figure 25-21: Distribution of Mine Operating Costs Excluding Power And Heating (AMC)

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25.7.7 Mineral Processing

25.7.7.1 Mineral Processing Total Operating Cost

Table 25-19 identifies the total mineral process operating cost for the New Afton Project.

Table 25-19: Total Mineral Process Unit Operating Cost(C$/tonne Milled)

Operation	Initial	Expanded	Life of mine
Labour	2.59	1.52	1.66
Consumable supplies	2.90	1.99	2.12
Maintenance supplies	0.53	0.39	0.41
Power	1.91	1.25	1.33
Total	7.93	5.15	5.51

25.7.8 General and Administration

The General and Administration compliment of the workforce has been estimated based on 26 employees initially, and 36 when operating at 4 million tonnes per year. The following table provides a summary of G&A labour, non labour and utilities costs.

Table 25-20: Summary of G&A Costs (C$/t milled)

Operation	Initial		Expanded	Life of mine
Labour	1.22		0.61	0.68
Non Labour	0.85	.	0.35	0.40
Power	0.33		0.25	0.27
Total	2.40		1.21	1.35

Concentrate transport costs are considered part of the net smelter return, calculated in the economic model and therefore were not listed in the summary of operating costs. The estimated cost of concentrate transport is given in Section 25.3

25.8 Economic Analysis (This section was written by John Shillabeer, P.Eng, Hatch)

The data generated in the study was used to construct an economic model of the project in order to assess the project’s economic parameters including the internal rate of return (IRR) and net present value (NPV).

The project economic model has been prepared based, in part, on the capital and operating costs estimates set out in this report. Therefore it should be assumed that the exclusions and assumptions that relate to the cost estimates also relate to the economic analysis (e.g. the occurrence of any of the risk factors identified in section 25.6.3.9 might have a material impact on the accuracy of this economic analysis).

25.8.1 Cash Flow Model

The economic model of the New Afton project includes the following assumptions:

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Table 25-21: Key Financial Assumptions

Financial Parameter	Assumption	Note
Currency	US$	The model results are quoted in US dollars
Discounted Cash Flow	16 years	The model DCF time frame covers the period from 2007 to 2022 inclusive.
(DCF) total time frame		Following a 2 year prime capital expenditure period for pre-production, there are
		2 years initial production at 1.6 Mt/ywhen expansion capital is also spent before
		ramping up to 4 Mt/y for the remaining life of the mine. The mine operating life is 13
		years . Closure costs are incurred in 2021 when there is some production and in
		2022. Working capital is returned in 2022.
DCF incremental time	Annual	A complete year is used as the DCF incremental time period in which all production
period	Discrete	costs and revenues are assumed as single point averages for the year. All cash flows
		applied to a particular year are assumed to occur as single points within that year.
Equity / debt financing	100% equity	The capital structure is estimated on a 100% equity basis, with no debt or interest
		payments. The cash flows are organized to represent the cash flows as seen by an
		equity investor owner.
Taxes	After tax	Calculated in accordance with prevailing rates and regulations (See section 25.5)
Integration	None	The project is assessed as an independent stand alone business for tax purposes.
WACC		Hatch did not estimate the weighted average cost of capital. Net present values were
		calculated using real discount rates of 0% and 5%
Inflation	Nil	The model is assessed in constant dollar terms. The base date of the model is 1
		January 07.
Terminal value	Nil	No terminal value is applied at the end of the project to account for the value of the
		ongoing operations (a conservative approach) but the working capital is presumed
		liquidated.

The key components of the model in creating the cash flow forecast are:

• summary of production volumes and contained grades of saleable elements

• metal prices

• revenue estimates from concentrate sales, commercial terms including TCs (treatment charges) and RCs (refining charges) and the payment terms for PMs (precious metal content: taken to be gold and silver only) and PP (copper price participation above the industry standard reference point)

• production parameters and expected metal recovery performance

• cash operating cost estimates for the mine and concentration facilities (excluding depreciation)

• capital cost and schedule including all associated indirect costs for project implementation

• working capital covering inventories, payable and receivables

• taxes

25.8.2 Metal Prices and Exchange Rate

The following long term, constant real terms metal prices and exchange rate were used in the economic model. Each metal price is the average closing price realised on the London Metal Exchange between January 2, 2004 and January 1, 2007 (three year trailing prices). The Canadian/US dollar exchange rate is also the average of the spot rates for the same period, as quoted by Moodys. Project results are quoted in US dollars corresponding to the international price quotations for metals. The economic analysis includes Canadian dollar amounts in the estimates of capital and operating costs, but these are then converted to US dollars to integrate the cash operating costs into the cash flow statement.

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Table 25-22: Metal Prices and Exchange Rate (3 Year Trailing)

Commodity	Units	Price
Copper	US$/lb Cu	2.01
Gold	US$/oz Au	487
Silver	US$/oz Ag	8.54
C$/US$ exchange rate		0.82

Hatch can offer no comment on future metal prices, exchange rates and inflation. Readers are encouraged to draw their own conclusions about the possible range of economic outcomes, assisted by the sensitivity charts presented in Figure 25-22 and Figures 25-23

25.8.3 Economic Results

Table 25-23: New Afton Project Summary of Economic Results

Copper price (LME)	US$ /lb	2.01
Gold price (LME)	US$ /oz	487
Silver price (LME)	US$ /oz	8.54
Exchange rate C$ / US$		0.82
After tax cash flow	US$ million	396.4
NPV (net present value) @ 0%before tax	US$ million	614.3
NPV (net present value) @ 5% before tax	US$ million	265.9
NPV (net present value) @ 5% after tax	US$ million	143.0
Equity IRR (internal rate of return before tax)	%	13.6
Equity IRR (internal rate of return after tax)	%	10.4
Cash cost (average life of mine)1	US$/lb Cu	0.64
Project payback (from 2009)	years	6.3

Note 1: net of by-product credits

Figure 25-22: Sensitivity Of IRR Before Tax To Variations In Costs And Metal Prices

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Figure 25-23: Sensitivity of Before Tax NPV @ 5% to Variations in Costs and Metal Prices

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26. Illustrations

Figure 26-1 shows the location of the proposed mine and surface plant relative to the existing Afton open pit and the Mineral Lease boundary.

Figure 26.2 shows the layout of the proposed plant area.

These illustrations are also available as separate PDF files, which can be printed in 11” x 17” format.

Illustrations
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