EX-99.64 3 a11-14376_2ex99d64.htm EX-99.64

Exhibit 99.64

TECHNICAL REPORT PREPARED FOR FRANCO-NEVADA CORPORATION IN RESPECT OF THE PROSPERITY GOLD-COPPER PROJECT, BRITISH COLUMBIA, CANADA, OWNED BY TASEKO MINES LIMITED

QUALIFIED PERSON:

SCOTT JONES, P.ENG.

Effective Date: May 12, 2010

Report Date: June 25, 2010

TABLE OF CONTENTS

1.		SUMMARY		7

2.		INTRODUCTION		12

3.		RELIANCE ON OTHER EXPERTS		14

4.		PROPERTY DESCRIPTION AND LOCATION		15

5.		ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY		19

	5.1		ACCESS	19
	5.2		RESOURCES AND INFRASTRUCTURE	20
	5.3		PHYSIOGRAPHY	21
	5.4		CLIMATE	21

6.		HISTORY		22

7.		GEOLOGICAL SETTING		25

8.		DEPOSIT TYPES		27

	8.1		SURFICIAL GEOLOGY	31
	8.2		VOLCANIC AND SEDIMENTARY ROCKS	32
	8.3		FISH LAKE INTRUSIVE COMPLEX	32
	8.4		ALTERATION	39
	8.5		STRUCTURE	39

9.		MINERALIZATION		40

10.		EXPLORATION		44

	10.1		EXTENT OF ALL RELEVANT EXPLORATION	44

11.		DRILLING		48

	11.1		DRILLING PRE-1991	48
	11.2		DRILLING 1991-1994	48
	11.3		DRILLING 1996-1997	48
	11.4		DRILLING 1998	49
	11.5		DRILLING 2007	49

12.		SAMPLING METHOD AND APPROACH		51

	12.1		CORE LOGGING	51
	12.2		SAMPLING	51

13.		SAMPLE PREPARATION, ANALYSES AND SECURITY		54

	13.1		SECURITY	54
	13.2		SAMPLE PREPARATION	54
	13.3		SAMPLE ANALYSIS 1991-2007	55
	13.4		QUALITY ASSURANCE QUALITY CONTROL	55
	13.5		SPECIFIC GRAVITY — BULK DENSITY MEASUREMENTS	60

14.		DATA VERIFICATION		61

	14.1		DATABASE	61
	14.2		VERIFICATION	62

15.		ADJACENT PROPERTIES		64

16.		MINERAL PROCESSING AND METALLURGICAL TESTING		65

	16.1		INTRODUCTION	65
	16.2		COMPOSITES	66
	16.3		MINERALOGY	69
	16.4		GRINDING	70
	16.5		GRAVITY SEPARATION	71
	16.6		BATCH TESTS	71
	16.7		LOCKED CYCLE TESTS	72
	16.8		RUN-IN PILOT PLANT AND MAIN PILOT PLANT RUNS	73
	16.9		TARGET CONCENTRATE GRADES AND RECOVERIES	74
	16.10		CONCENTRATE ANALYSIS	77
	16.11		TAILINGS SETTLING TESTS	78
	16.12		ENVIRONMENTAL DATA	79
	16.13		REGRIND EVALUATION TEST WORK	79

17.		MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES		80

	17.1		RESOURCE MODELING	80
	17.2		RESOURCE CLASSIFICATION	81
	17.3		RESERVE ESTIMATION	83

18.		ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES
AND PRODUCTION PROPERTIES				89

	18.1		SITE INFRASTRUCTURE	89
	18.2		OPEN PIT DESIGN	102
	18.3		MINING OPERATIONS	111
	18.4		PROCESSING AND CONCENTRATOR	117
	18.6		RECOVERABILITY	124
	18.7		MARKETS	124
	18.8		CONTRACTS	127
	18.9		ENVIRONMENTAL CONSIDERATIONS	127
	18.10		TAXES	132
	18.11		CAPITAL AND OPERATING COST ESTIMATES	132
	18.12		ECONOMIC ANALYSIS	147

19.		INTERPRETATIONS AND CONCLUSIONS		157

20.		RECOMMENDATIONS		160

21.		REFERENCES		161

22.		DATE AND SIGNATURE PAGE		164

LIST OF FIGURES

FIGURE 4-1	PROSPERITY LOCATION	15
FIGURE 4-2	CLAIM MAP	16
FIGURE 5-1	PROSPERITY LOCATION	20
FIGURE 7-1	REGIONAL GEOLOGY	26
FIGURE 8-1	GEOLOGY AT OVERBURDEN-BEDROCK INTERFACE	28
FIGURE 8-2	SURFICIAL GEOLOGY	30
FIGURE 8-3	GEOLOGY PLAN VIEW 1400M ELEVATION	34
FIGURE 8-4	GEOLOGY PLAN VIEW 1200M ELEVATION	35
FIGURE 8-5	GEOLOGY PLAN VIEW 1000M ELEVATION	36
FIGURE 8-6	GEOLOGY VERTICAL SECTION 10100E	37
FIGURE 8-7	GEOLOGY VERTICAL SECTION 10100N	38
FIGURE 9-1	AU AND CU GRADES AT 1400M ELEVATION	41
FIGURE 9-2	AU AND CU GRADES AT 1200M ELEVATION	42
FIGURE 9-3	AU AND CU GRADES AT 1000M ELEVATION	43
FIGURE 10-1	DRILL HOLE LOCATIONS 1969 TO 1994	46
FIGURE 10-2	DRILL HOLE LOCATIONS 1996 TO 1998	47
FIGURE 12-1	1996-1997 DRILL CORE SAMPLING, PREPARATION & ANALYTICAL FLOW CHART	53
FIGURE 16-1	PHASE II TEST PROGRAM — DRILL HOLE PLAN	66
FIGURE 16-2	PHASE III TEST PROGRAM — DRILL HOLE PLAN	67
FIGURE 16-3	PHASE IV TEST PROGRAM — DRILL HOLE PLAN	68
FIGURE 16-4	TARGET COPPER RECOVERY VS COPPER HEAD GRADE	76
FIGURE 16-5	TARGET CONCENTRATE COPPER GRADE VS COPPER HEAD GRADE	76
FIGURE 16-6	TARGET GOLD RECOVERY VS GOLD HEAD GRADE	77
FIGURE 17-1	GEMCOM-WHITTLE PIT BY PIT GRAPH	85
FIGURE 18-1	GENERAL SITE LAYOUT	89
FIGURE 18-2	PLANT LAYOUT	93
FIGURE 18-3	TAILINGS CONTAINMENT	97
FIGURE 18-4	MAIN EMBANKMENT	100
FIGURE 18-5	WEST EMBANKMENT	101
FIGURE 18-6	GEOTECHNICAL PIT SLOPE DESIGN SECTORS PLAN	105
FIGURE 18-7	PROSPERITY FINAL PIT — END OF YEAR 31	112
FIGURE 18-8	MATERIAL MOVEMENT SCHEDULE	115
FIGURE 18-9	SIMPLIFIED FLOWSHEET	118
FIGURE 18-10	CONSTANT $ COPPER CORRELATION WITH CDN/US CURRENCY EXCHANGE	148
FIGURE 18-11	HISTORICAL CORRELATION — COPPER AND EXCHANGE RATE 1998 TO 2006	149
FIGURE 18-12	ROI SENSITIVITY	156

LIST OF TABLES

TABLE 1-1	PROSPERITY MINERAL RESERVES	7
TABLE 1-2	PROSPERITY MINERAL RESOURCES	7
TABLE 4-1	PROSPERITY MINERAL CLAIMS	17
TABLE 6-1	DRILLING SUMMARY 1963-1989	23
TABLE 8-1	PROSPERITY GOLD-COPPER PROJECT GEOLOGY CODES	29
TABLE 10-1	DRILLING SUMMARY 1963-1998	45
TABLE 11-1	DRILL HOLES BY ORIENTATION BY YEAR	49
TABLE 12-1	NUMBER OF SAMPLES BY YEAR	52
TABLE 13-1	QAQC SAMPLE TYPES USED	56
TABLE 13-2	DRILL HOLE SAMPLE QAQC SUMMARY	57
TABLE 13-3	SUMMARY OF COPPER-GOLD STANDARD REFERENCE MATERIALS USED	58
TABLE 13-4	SPECIFIC GRAVITY MEASUREMENTS BY YEAR AND METHOD	60
TABLE 16-1	METRIC WORK INDICES	70
TABLE 16-2	TARGET GOLD & COPPER RECOVERIES AND CONCENTRATE GRADES	74
TABLE 16-3	TARGET COPPER RECOVERY & TARGET CONCENTRATE COPPER GRADE CALCULATIONS: LOWER, MIDDLE AND UPPER ZONES	74
TABLE 16-4	TARGET GOLD RECOVERY CALCULATIONS UPPER ZONE	75
TABLE 16-5	TARGET GOLD RECOVERY CALCULATIONS MIDDLE & LOWER ZONES	75
TABLE 16-6	TYPICAL CONCENTRATE ANALYSIS	77
TABLE 16-7	SETTLING TESTS ON CONCENTRATE FROM UPPER COMPOSITE	78
TABLE 17-1	BLOCK MODEL EXTENTS	80
TABLE 17-2	PROSPERITY MINERAL RESOURCES	83
TABLE 17-3	PIT OPTIMIZATION PARAMETERS	84
TABLE 17-4	NSR PIT RIM CUT-OFF CALCULATION	86
TABLE 17-5	2009 NET SMELTER RETURN CALCULATIONS	87
TABLE 18-1	ELECTRICAL LOAD ANALYSIS — YEAR 7	92
TABLE 18-2	EMBANKMENT CONSTRUCTION MATERIAL REQUIREMENTS	99
TABLE 18-3	RECOMMENDED WALL SLOPES	106
TABLE 18-4	DESIGN PARAMETERS	109
TABLE 18-5	MINE PRODUCTION FORECAST	114
TABLE 18-6	PREDICTED RECOVERIES & GRADES	120
TABLE 18-7	PRE-PRODUCTION CAPITAL COST ($ X 1000)	133
TABLE 18-8	PRE-PRODUCTION INDIRECT COSTS ($ X 1000)	134
TABLE 18-9	SUSTAINING CAPITAL COST ($ X 1000)	134
TABLE 18-10	LIFE-OF-MINE UNIT COSTS	140
TABLE 18-11	SUMMARY OF ESTIMATED AVERAGE ANNUAL OPERATING MANPOWER (YEARS 10-25)	142
TABLE 18-12	LIFE-OF-MINE DIRECT MINING UNIT COST	143
TABLE 18-13	TYPICAL ANNUAL PROCESSING OPERATING COST SUMMARY (YEAR 7-25)	144
TABLE 18-14	ESTIMATED TYPICAL ANNUAL GENERAL & ADMINISTRATION COSTS (YEAR 10-25) ($000’S)	146
TABLE 18-15	KEY RESERVE INDICATORS	147
TABLE 18-16	NSR ASSUMPTIONS	150
TABLE 18-17	OPERATING UNIT COSTS	151
TABLE 18-18	CAPITAL COST	152
TABLE 18-19	RESERVE BASIS CASH FLOW	154
TABLE 19-1	PROSPERITY MINERAL RESOURCES	158
TABLE 19-2	PROSPERITY MINERAL RESERVES	159

Abbreviation		Unit or Description
Ag		silver
amsl		above mean sea level
Au		gold
B.C.		British Columbia, Canada
BCEA		British Columbia Environmental Assessment Act
BE		Break Even
BH		Brook Hunt and Associates
BME		Bloomsbury Mineral Economics
ERA		Environmental Risk Assessment
CDN$		Canadian Dollars
CEAA		Canadian Environmental Assessment Act
CRU		Copper Research Unit
Cu		copper
DFO		Department of Fisheries
FLAC		Computer model
G&A		General and Administration
GCL		Giroux Consultants Ltd.
gpt		grams per tonne
Gwh		Gigawatt-hour
ha		hectare
IRR		internal rate of return
km		kilometre
kV		kilovolt
KP		Knight Piesold Consulting
lb		pound (weight)
m		metre
MIBC		Collector Reagent
mPa		megaPascal
Mt		million tonnes
µm		micron
NI		National Instrument 43-101
NMS		Nilsson Mine Services
NPV		net present value
NSR		net smelter return
NTS		National Topographic System
oz		Troy ounce
%		percent
PAG		Potentially Acid Generating
RA		Responsable Agency
RMR		Rock Mass Rating
SAG		Semi Autogenous Grinding
SG		specific gravity
SIBX		Collector Reagent
stpd		short tons per day
t		tonne (metric)
US$		United States Dollars
TC		Treatment Charge
TCRC		Treatment and Refining Charge
TCRCPP		Treatment, Refining and Price Participation
TWC 314		Mill Reagent
TWC 401		Mill Reagent

1. Summary

Taseko Mines Limited (“Taseko” or the “Company”) has increased the mineral reserve by 344 million tonnes at its 100% owned Prosperity gold-copper project (the “Prosperity Project”), indicating that the property hosts proven and probable reserves of 831 million tonnes grading 0.41 gpt Au and 0.23% Cu.

The reserve was previously based on metal prices of US$1.50/lb for copper and US$500/oz for gold, an exchange rate of US$0.80/CDN$ 1.00, and a $5.25 net smelter return per tonne (NSR/t) cut-off.

Updated reserves are based on copper and gold prices of US$1.65/lb and US$650/oz respectively, an exchange rate of US$0.82/CDN$ 1.00, and a $5.50/t NSR cut-off

The mineral reserves are estimated as shown in Table 1-1.

Table 1-1

Prosperity Mineral Reserves

at CDN$5.50 NSR/t Pit-Rim Cut-off
				Recoverable	Recoverable
	Tonnes	Gold	Copper	Gold Ounces	Copper Pounds
Category	(millions)	(gpt)	(%)	(millions)	(billions)
Proven	481	0.46	0.26	5.0	2.4
Probable	350	0.35	0.18	2.7	1.2
Total	831	0.41	0.23	7.7	3.6

The reserve estimate takes into consideration all geologic, mining, milling, and economic factors, and is stated according to Canadian standards (NI43-101).

The Proven and Probable Reserves are included in the Measured and Indicated Mineral Resources shown in Table 1-2. The Mineral Resources are estimated at a 0.14% Cu cut-off.

Table 1-2

Prosperity Mineral Resources

at 0.14% Copper Cut-off
	Tonnes	Gold	Copper
Category	(millions)	(gpt)	(%)
Measured	547.1	0.46	0.27
Indicated	463.4	0.34	0.21
Total	1,010.5	0.41	0.24

The Prosperity Project is located 125 km southwest of the City of Williams Lake in the Cariboo-Chilcotin region of British Columbia, Canada.

The mining claims are 100% owned by Taseko, are not subject to any royalties or carried interests and are currently in good standing until the year 2018. The claims are situated within an area that was the subject of an aboriginal Rights and Title case in which the Supreme Court of B.C. recently found that the Tsilhqot’in Nation’s rights do not include the subsurface rights to the area around Fish Lake.

Taseko carried out ongoing and systematic exploration programs on the Project from 1991 — 1999, increasing drilling to 156,339 m in 470 holes, outlining a large porphyry gold-copper deposit. A comprehensive audit and verification program of the geology and assay results in 1998 found the geological work for the Prosperity to be done in a professional manner and according to industry standards.

The Company and its consultants also carried out progressive engineering, metallurgical and environmental studies over the period 1998 to 2009 including a feasibility level study of the Prosperity Project in 2000 by Kilborn SNC Lavalin and a mill redesign and project cost review in 2006 by SNC Lavalin in 2006. Taseko utilized information from the 2000 feasibility level study and the 2006 revised process design to prepare a pre-feasibility study, the results of which were announced in a Taseko News Release dated January 11, 2007 and summarized in the 43-101 Technical Report dated February 25, 2007.

Consistent with the recommendations of the February 25, 2007 Technical Report, HATCH, Knight Piesold Consulting, and Taseko Mines Limited completed a feasibility study update incorporating the 2000 SNC Feasibility Study, 2006 SNC Lavalin Mill Redesign, additional revisions to the processing plant and infrastructure, updates to the tailings facility design and pit geotechnical analysis, and revisions to the design and scheduling of the open pit. The results were announced in a Taseko News Release dated September 24, 2007 and summarized in the 43-101 Technical Report dated October 15, 2007.

In 2008 Taseko worked with Axxent Engineering Ltd and Rutter Hinz to investigate value engineering opportunities, energy efficiency, and operating ease in various areas of the concentrator and support infrastructure.

In 2009 a significantly different outlook for copper and gold prices from that of 2007 warranted a re-evaluation of the reserve.

A technical report dated December 17, 2009 (with an effective date of November 2, 2009) (the “Initial report”) was prepared by Taseko to document the results of the evaluation of increased metal prices on the reserves at Prosperity as announced in a Taseko News Release dated November 2, 2009 in the format prescribed in National Instrument 43-101. The Initial Report has been readdressed to Franco-Nevada Corporation (“Franco-Nevada”) in connection with its press release dated May 12, 2010 announcing the execution by Franco-Nevada and Taseko of a purchase and sale agreement (the “Gold Stream Agreement”) relating to the purchase by Franco-

Nevada of gold produced at the Prosperity project. Although the Initial Report has been updated to the effective date of May 12, 2010, there are no material changes.

The proposed mine plan utilizes a large-scale conventional truck shovel open pit mining and milling operation. Following a one and a half year pre-strip period, total material mined from the open pit over years 1 through 31 averages 170,000 tonnes/day at a life of mine strip ratio of 1.5:1. A declining net smelter return cut-off is applied to the mill feed, which defers lower grade ore for later processing. The stockpiled ore is processed in the final years of the mine plan.

The Prosperity processing plant has been designed with a nominal capacity of 70,000 tonnes/day. The plant consists of a single 12-m diameter semi-autogenous grinding (SAG) mill, two 7.9-m diameter ball mills, followed by processing steps that include bulk rougher flotation, regrinding, cleaner flotation, thickening and filtering to produce a copper-gold concentrate. Expected life-of-mine metallurgical recovery is 87% for copper and 69% for gold, with annual production averaging 110 million pounds copper and 234,000 ounces gold over the 33 year mine life.

The copper-gold concentrate will be hauled with highway trucks to an expanded load-out facility at the Gibraltar Mines Ltd.’s existing facility near Macalister for rail transport to various points of sale, but mostly through the Port of Vancouver for shipment to smelters/refineries around the world.

Power will be supplied via a new 124 km long, 230 kV transmission line from Dog Creek on the BC Hydro Grid. Infrastructure would also include the upgrade of sections of the existing road to the site, construction of a short spur to the minesite, an on-site camp, equipment maintenance shop, administration office, concentrator facility, warehouse, and explosives facilities.

Based on this update, the project would employ up to 460 permanent hourly and staff personnel. In addition, approximately 60 contractor personnel would be employed in areas including catering, concentrate haulage, explosives delivery, and bussing.

Key project metrics include:

· 3.6B pounds of recoverable copper

· 7.7M ounces of recoverable gold

· 33 year mine life at a milling rate of 70,000 tonnes/day

· Life of mine waste to ore strip ratio of 1.5

· Total pre-production capital cost of CDN$814 million

· Site operating cost of CDN$7.51 per tonne milled over the life of mine

· Total operating costs net of byproduct credits of US$0.59/lb Cu

The project is currently in the harmonized British Columbia Environmental Assessment Act (EA Act) and Canadian Environmental Assessment Act (CEAA) review process. Decisions by provincial Ministers as to the issuance of a provincial Environmental Assessment Certificate are anticipated in early 2010. Decisions by the federal cabinet as to the issuance of a federal Environmental Assessment Certificate are anticipated in mid-2010. If approved the Environmental Assessment Certificates will be for the 487M tonne reserve, 70,000 tonnes per

day mine plan outlined in the Environmental Impacts Assessment Report and not the 830M tonne reserve announced on November 2nd 2009 that is the subject of this report.

Conclusion and Recommendation

The project is technically and economically viable under the assumptions of the 2009 reserve update.

The drilling, geological interpretation and resource modeling have been done in a professional manner and to industry standards.

The project proposes to utilize mining equipment and operating practices that are industry standard. The open pit design is based on sound geotechnical investigations, recommendations and design criteria, and incorporates adequate stability design and dewatering parameters

The Prosperity concentrator design has been based on suitable metallurgical test work and incorporates proven technologies and equipment.

The tailings storage facility has been designed incorporating adequate site characteristics, geotechnical, hydrogeological, and water management considerations for the purposes of the study.

The capital cost estimate reflects the design changes resulting from the 2008 engineering. Both the capital and operating cost estimates are considered to reflect 2nd quarter 2009.

In the opinion of the author, the geological interpretation, resource model, mine plan, metallurgical test work, concentrator design and supporting infrastructure are suitable for this reserve estimate.

The following are the principal risk factors and uncertainties which, in the author’s opinion, are likely to most directly affect the ultimate feasibility of the Prosperity project. The mineralized material at the Prosperity project is currently classified as a measured and indicated resources, and a portion of it qualifies under Canadian mining disclosure standards as a proven and probable reserve, but readers are cautioned that no part of the Prosperity project’s mineralization is yet considered to be a reserve under US mining standards as all necessary mining permits would be required in order to classify the project’s mineralized material as an economically exploitable reserve.

Although feasibility level work has been done to confirm the mine design, mining methods and processing methods assumed in the reserve update, construction and operation of the mine and processing facilities depend on securing environmental and other permits on a timely basis.

Additional permits, when required, have yet to be applied for and there can be no assurance that required permits can be secured or secured on a timely basis or that third party opposition will not exist, which may delay or otherwise affect the Company’s ability to secure required permits.

Although costs, including design, procurement, construction and on-going operating costs and metal recoveries have been established at an appropriate level of detail required for supporting a reserve estimate, these could be materially different from those experienced during actual execution and operation. There can be no assurance that these infrastructure facilities can be developed on a timely and cost-effective basis. Energy risks include the potential for significant increases in the cost of fuel and electricity. The reserve update assumes specified, long-term prices levels for gold and copper. The prices of these metals have historically been volatile, and the Company has no control of or influence on the prices, which are determined in international markets. There can be no assurance that the price of gold and copper will continue at current levels or that these prices will not decline below the prices assumed in the reserve update. Prices for gold and copper have been below the price ranges assumed in reserve update at times during the past ten years, and for extended periods of time. The project will require major financing, probably a combination of debt and equity financing. Although interest rates are at historically low levels, there can be no assurance that debt and/or equity financing will be available on acceptable terms.

Other general risks include those typical of very large construction projects, including the general uncertainties inherent in engineering and construction cost, the need to comply with generally increasing environmental obligations, and accommodation of local and community concerns.

2. Introduction

The Qualified Persons responsible for the content of this report are:

Scott Jones, P.Eng., Vice President of Engineering for Taseko. Mr. Jones has reviewed the methods used to determine grade and tonnage in the geological model, reviewed the long range mine plan, the capital and operating cost estimates, and directed the updated economic evaluation. He has visited the Prosperity property on numerous occasions from 2006 to 2009.

G.H. Giroux, P.Eng., MASc., independent consulting geological engineer. Mr. Giroux was responsible for the Resource Estimation Section completed in Vancouver during 1998 and amended 1999. He has not visited the property.

Lawrence Melis, P.Eng., consulting process engineer, working for Melis Engineering Ltd. Mr. Melis was responsible for the metallurgical test work completed in the 1990’s by Melis Engineering Ltd. He visited the property in the 1990’s to look at core and general site conditions.

The following information has been relied upon as provided by qualified persons who have provided certificates in Section 22:

Giroux, G.H., 1998. A Resource Estimate Update for the Prosperity Project Gold-Copper Deposit. Unpublished Company Report. Taseko Mines Limited, Vancouver, British Columbia.

Giroux, G.H., 1999. Addendum to the March 12, 1998 Resource Estimate for the Prosperity Project Gold-Copper Deposit for Taseko Mines Limited. Unpublished Company Report, Taseko Mines Limited, Vancouver, British Columbia.

Melis Engineering Ltd., 1998. Prosperity Gold-Copper Project Feasibility Study, Volume 4, Appendix E, Metallurgy.

Additional sources of information used for this report are:

2008 Engineering by Axxent Engineering Ltd under the supervision of Mark Dobbs, P.Eng.

2008 Engineering by Rutter Hinz under the supervision of Ray Turenne, P.Eng.

2008 Engineering by Farnell Thompson under the supervision of Doug Farnell, P.Eng.

2008 Engineering by HATCH Energy under the supervision of Robert Henderson, P.Eng..

2007 Prosperity Feasibility Study by HATCH under the supervision of Steve McMaster, P.Eng

2007 to 2009 tailings, water balance, and geotechnical studies conducted by Knight Piesold Ltd., under the supervision of Ken Brouwer, P.Eng.

Sampling, Analysis and Quality Assurance/Quality Control by Eric Titley, P.Geo.

Mineral reserves, mine planning and design aspects were developed by AKF Mining Services Inc. under the supervision of Taseko Mines Limited.

All of the above persons are independent of the Company except for Mr. Titley and Mr. Jones.

3. Reliance on Other Experts

In preparing this Technical Report, the author relied on the following information which may not be by qualified persons:

Information on History, Property, Deposit, and Mineralization was acquired from Taseko Mines Limited.

The author is not an expert on mineral tenure and has depended on the information received from the Company.

4. Property Description and Location

The Prosperity property is located approximately 125 km southwest of the City of Williams Lake in south-central British Columbia, Canada (Figure 4-1) at latitude 51 degrees 28’ N and longitude 123 degrees 37’ W.

The property is located in the Clinton Mining Division on the N.T.S. map sheet 92 O/SE. The 85 square km property is comprised of 124 mineral claims depicted in Figure 4-2. The claims are 100% owned by Taseko and are not subject to any royalties or carried interests.

The mineral claims are currently in good standing until the year 2018 as shown in Table 4-1.

Table 4-1

Prosperity Mineral Claims

Taseko Mines Limited

(Corporate Free Miner 126450)

Prosperity Project, Clinton Mining Division

Mineral Claims

Tenure Number	Claim Name	Owner	Map Number	Good To Date	Status	Area (ha)

208019	BCC #5FR.	126450 (100%)	092O	2018/apr/02	GOOD	25
208020	BCC #6FR.	126450 (100%)	092O	2018/apr/02	GOOD	25
208024	EKO 1	126450 (100%)	092O	2018/apr/02	GOOD	500
208025	EKO 2	126450 (100%)	092O	2018/apr/02	GOOD	500
208026	EKO 3	126450 (100%)	092O	2018/apr/02	GOOD	500
209279	TKO 2	126450 (100%)	092O	2018/apr/02	GOOD	500
209324	FISH 1	126450 (100%)	092O	2018/apr/02	GOOD	500
209325	FISH 2	126450 (100%)	092O	2018/apr/02	GOOD	500
209326	FISH 3	126450 (100%)	092O	2018/apr/02	GOOD	500
209327	FISH 4	126450 (100%)	092O	2018/apr/02	GOOD	500
209487	BJ #1	126450 (100%)	092O	2018/apr/02	GOOD	25
209488	BJ #3	126450 (100%)	092O	2018/apr/02	GOOD	25
209489	BJ #5	126450 (100%)	092O	2018/apr/02	GOOD	25
209490	BJ #7	126450 (100%)	092O	2018/apr/02	GOOD	25
209491	BJ #9	126450 (100%)	092O	2018/apr/02	GOOD	25
209492	BJ #11	126450 (100%)	092O	2018/apr/02	GOOD	25
209496	BJ #16	126450 (100%)	092O	2018/apr/02	GOOD	25
209497	BJ #17	126450 (100%)	092O	2018/apr/02	GOOD	25
209498	BJ #18	126450 (100%)	092O	2018/apr/02	GOOD	25
209499	BJ #19	126450 (100%)	092O	2018/apr/02	GOOD	25
209500	BJ #20	126450 (100%)	092O	2018/apr/02	GOOD	25
209501	BJ #21	126450 (100%)	092O	2018/apr/02	GOOD	25
209502	BJ #22	126450 (100%)	092O	2018/apr/02	GOOD	25
209503	BJ #23	126450 (100%)	092O	2018/apr/02	GOOD	25
209504	BJ #24	126450 (100%)	092O	2018/apr/02	GOOD	25
209509	BJ #29	126450 (100%)	092O	2018/apr/02	GOOD	25
209511	BJ #31	126450 (100%)	092O	2018/apr/02	GOOD	25
209512	BJ #32	126450 (100%)	092O	2018/apr/02	GOOD	25
209513	BJ #33	126450 (100%)	092O	2018/apr/02	GOOD	25
209514	BJ #34	126450 (100%)	092O	2018/apr/02	GOOD	25
209515	BJ #35	126450 (100%)	092O	2018/apr/02	GOOD	25
209516	BJ #36	126450 (100%)	092O	2018/apr/02	GOOD	25
209517	BJ #37	126450 (100%)	092O	2018/apr/02	GOOD	25
209519	BJ #39	126450 (100%)	092O	2018/apr/02	GOOD	25
209520	BJ #40	126450 (100%)	092O	2018/apr/02	GOOD	25
209521	BJ #41	126450 (100%)	092O	2018/apr/02	GOOD	25
209522	BJ #42	126450 (100%)	092O	2018/apr/02	GOOD	25
209535	L7	126450 (100%)	092O	2018/apr/02	GOOD	25
209536	L8	126450 (100%)	092O	2018/apr/02	GOOD	25
209537	L9	126450 (100%)	092O	2018/apr/02	GOOD	25
209538	L10	126450 (100%)	092O	2018/apr/02	GOOD	25
209539	L11	126450 (100%)	092O	2018/apr/02	GOOD	25
209540	L12	126450 (100%)	092O	2018/apr/02	GOOD	25
209541	L21	126450 (100%)	092O	2018/apr/02	GOOD	25
209542	L22	126450 (100%)	092O	2018/apr/02	GOOD	25
209543	L23	126450 (100%)	092O	2018/apr/02	GOOD	25
209544	L24	126450 (100%)	092O	2018/apr/02	GOOD	25
209545	L31	126450 (100%)	092O	2018/apr/02	GOOD	25
209546	L32	126450 (100%)	092O	2018/apr/02	GOOD	25
209547	L33	126450 (100%)	092O	2018/apr/02	GOOD	25
209548	L34	126450 (100%)	092O	2018/apr/02	GOOD	25
209549	L35	126450 (100%)	092O	2018/apr/02	GOOD	25
209550	L36	126450 (100%)	092O	2018/apr/02	GOOD	25
209551	L37	126450 (100%)	092O	2018/apr/02	GOOD	25
209552	L38	126450 (100%)	092O	2018/apr/02	GOOD	25
209553	L39	126450 (100%)	092O	2018/apr/02	GOOD	25
209554	L40	126450 (100%)	092O	2018/apr/02	GOOD	25
209555	L41	126450 (100%)	092O	2018/apr/02	GOOD	25
209556	L42	126450 (100%)	092O	2018/apr/02	GOOD	25
209557	L43	126450 (100%)	092O	2018/apr/02	GOOD	25
209558	L44	126450 (100%)	092O	2018/apr/02	GOOD	25
209559	L45	126450 (100%)	092O	2018/apr/02	GOOD	25
209560	L46	126450 (100%)	092O	2018/apr/02	GOOD	25
209561	L47	126450 (100%)	092O	2018/apr/02	GOOD	25
209562	L48	126450 (100%)	092O	2018/apr/02	GOOD	25
209572	K66	126450 (100%)	092O	2018/apr/02	GOOD	25
209578	K116	126450 (100%)	092O	2018/apr/02	GOOD	25
209579	K117	126450 (100%)	092O	2018/apr/02	GOOD	25
209580	K118	126450 (100%)	092O	2018/apr/02	GOOD	25
209581	K119	126450 (100%)	092O	2018/apr/02	GOOD	25
209582	K120	126450 (100%)	092O	2018/apr/02	GOOD	25
209583	K121	126450 (100%)	092O	2018/apr/02	GOOD	25
209584	K125	126450 (100%)	092O	2018/apr/02	GOOD	25
209585	K126	126450 (100%)	092O	2018/apr/02	GOOD	25
209586	K127	126450 (100%)	092O	2018/apr/02	GOOD	25
209587	K128	126450 (100%)	092O	2018/apr/02	GOOD	25
209588	K129	126450 (100%)	092O	2018/apr/02	GOOD	25
209589	K130	126450 (100%)	092O	2018/apr/02	GOOD	25
209590	K131	126450 (100%)	092O	2018/apr/02	GOOD	25
209591	K132	126450 (100%)	092O	2018/apr/02	GOOD	25
209592	K133	126450 (100%)	092O	2018/apr/02	GOOD	25
209593	K134	126450 (100%)	092O	2018/apr/02	GOOD	25
209594	K135	126450 (100%)	092O	2018/apr/02	GOOD	25
209595	K136	126450 (100%)	092O	2018/apr/02	GOOD	25
209598	TEL #75	126450 (100%)	092O	2018/apr/02	GOOD	25
209611	TK #15	126450 (100%)	092O	2018/apr/02	GOOD	25
209619	TK #23	126450 (100%)	092O	2018/apr/02	GOOD	25
209621	TK #25	126450 (100%)	092O	2018/apr/02	GOOD	25
209622	TK #26	126450 (100%)	092O	2018/apr/02	GOOD	25
209640	TK #46	126450 (100%)	092O	2018/apr/02	GOOD	25
209656	TK #67	126450 (100%)	092O	2018/apr/02	GOOD	25
314004	F 2	126450 (100%)	092O	2018/apr/02	GOOD	25
314005	F 3	126450 (100%)	092O	2018/apr/02	GOOD	25
314006	F 4	126450 (100%)	092O	2018/apr/02	GOOD	25
314007	F 5	126450 (100%)	092O	2018/apr/02	GOOD	25
314008	F 6	126450 (100%)	092O	2018/apr/02	GOOD	25
314009	F 7	126450 (100%)	092O	2018/apr/02	GOOD	25
314010	F 8	126450 (100%)	092O	2018/apr/02	GOOD	25
314025	F 9	126450 (100%)	092O	2018/apr/02	GOOD	25
314026	FISH 10	126450 (100%)	092O	2018/apr/02	GOOD	300
314028	FISH 6	126450 (100%)	092O	2018/apr/02	GOOD	500
314029	FISH 7	126450 (100%)	092O	2018/apr/02	GOOD	500
314031	FISH 9	126450 (100%)	092O	2018/apr/02	GOOD	200
516779		126450 (100%)	092O	2018/apr/02	GOOD	504
516785		126450 (100%)	092O	2018/apr/02	GOOD	504
516849		126450 (100%)	092O	2018/apr/02	GOOD	625
516915		126450 (100%)	092O	2018/apr/02	GOOD	81
516926		126450 (100%)	092O	2018/apr/02	GOOD	1047
516935		126450 (100%)	092O	2018/apr/02	GOOD	363
516970		126450 (100%)	092O	2018/apr/02	GOOD	221
516984		126450 (100%)	092O	2018/apr/02	GOOD	60
516990		126450 (100%)	092O	2018/apr/02	GOOD	262
517288		126450 (100%)	092O	2018/apr/02	GOOD	81
517338		126450 (100%)	092O	2018/apr/02	GOOD	161
517347		126450 (100%)	092O	2018/apr/02	GOOD	20
517352		126450 (100%)	092O	2018/apr/02	GOOD	181
537996		126450 (100%)	092O	2018/apr/02	GOOD	20
537997		126450 (100%)	092O	2018/apr/02	GOOD	20
593521	PROS FR 1	126450 (100%)	092O	2018/apr/02	GOOD	20
593522	PROS FR 2	126450 (100%)	092O	2018/apr/02	GOOD	20
593523	PROS FR 3	126450 (100%)	092O	2018/apr/02	GOOD	40
593524	PROS FR 4	126450 (100%)	092O	2018/apr/02	GOOD	20
593525	PROS FR 5	126450 (100%)	092O	2018/apr/02	GOOD	20
593655	PROS FR 6	126450 (100%)	092O	2018/apr/02	GOOD	60

In the late 1990’s, Kilborn SNC Lavalin undertook a spot check of claim posts on the property during the drilling verification program. The spot checks concentrated on verifying a number of Legal Corner Posts and Identification Posts. Based on site inspections combined with examination of the Mineral Titles maps and documents as well as discussions with the B.C. Mineral Titles Branch staff in Vancouver, it was Kilborn’s opinion that the Taseko claim title to the project area was secure and legal.

In 2005, Taseko converted several of the ground staked legacy claims covering the mineralized area to cell claims as allowed by the amended BC Mineral Tenure Act. The legacy claim conversion consolidated the project holdings and eliminated any internal claim gaps.

In 2008 a legal survey of the area outlined in Figure 4-2 was conducted by Rathbone & Goodrich, BC & Canada Land Surveyors. Application has been made to convert the status of this area from mineral claims to a mining lease.

As this is a new project, the only environmental liabilities on the property are those related to remaining reclamation of exploration activities which are fully funded in the form of a reclamation deposit held by a financial institution on behalf of the B.C. Ministry of Finance

The Company does not hold any surface rights.

The Prosperity deposit is situated within an area that was the subject of an aboriginal Rights and Title case; Tsilhqot’in Nation v. British Columbia. A recent Supreme Court of B.C. ruling found that 100% of the areas that the Tsilhqot’in Nation claimed title to were in fact subject to aboriginal rights. The Tsilhqot’in Nation’s rights have been determined by the Court not to include the subsurface rights to the area around Fish Lake.

On January 6, 2009 Marilyn Baptiste on her own behalf and on behalf of the Xeni Gwet’in First Nation filed a statement of claim in the Supreme Court of British Columbia. The Plaintiff seeks an injunction to prevent the provincial and federal governments from issuing any permits required to advance the development of the Prosperity project. The matter is not being advanced by the plaintiff at this time.

Pursuant to the Gold Stream Agreement, Franco-Nevada agreed to acquire from Taseko an amount of gold equal to 22% of the gold produced at the Prosperity project. Franco-Nevada will provide a US$350 million deposit for the construction of the Prosperity project advanced pro rata with other financing for the project once the project is fully permitted and financed and will issue to Taseko with the first installment share purchase warrants exercisable to purchase 2,000,000 common shares of Franco-Nevada at $75.00 per share until June 2017. In addition, Franco-Nevada will pay Taseko the lower of US$400 an ounce (subject to an adjustment for inflation) or the prevailing market price for each ounce of gold delivered.

5. Accessibility, Climate, Local Resources, Infrastructure, and Physiography

5.1 Access

Current vehicle access to the site from Williams Lake is via Provincial Highway No. 20 to Lee’s Corner and forestry resource roadways (Figure 5-1). The access between Williams Lake and the mine site (approximately 180 km) must be an all-weather road and will comprise a portion of the following:

· Provincial Highway No. 20 - 90 km of 2 lane, paved road;

· Taseko Lake (Whitewater) Logging Road - 68 km of one-lane, 5 m wide, gravel road with turnouts;

· 4500 Road (Riverside Haul Road) - 19 km of one-lane, 5 m wide, gravel road with turn-outs;

· and a new Project Site Access Road - 2.8 km of 2 lane, 8 m wide, gravel road.

This road system will serve as the principal access road both during construction and mine operation.

5.2 Resources and Infrastructure

The City of Williams Lake supplies goods and services to two operating mines in the area; Imperial Metals’ Mount Polley Mine and Taseko’s Gibraltar Mine.

Multiple high-voltage transmission lines from the existing Peace River hydroelectric power grid are situated 118 km east of the Prosperity Project. A 124 km conventional power line designed to connect to the existing BC Hydro electric power grid will be capable of supplying the required power to service a large mine and mill complex at the Prosperity Project site. BC Hydro has confirmed, through an Interconnection System Impact Study completed by SNC Lavalin in June, 2007, and a Definition Stage Study along with Preliminary Engineering and Estimate Study in 2008 that the supply of power to the Prosperity project is technically viable through a proposed switching station at Dog Creek.

Sufficient water is available on the property for a mining operation.

The Canadian National Railway services Williams Lake and Gibraltar’s existing concentrate load-out facilities near Macalister, just north of McLeese Lake, and has the ability to move copper concentrates through to the Pacific Ocean Port of Vancouver and to smelter facilities and ports in eastern North America.

5.3 Physiography

The property is located on the Fraser Plateau in the Taseko Lakes region on the eastern side of the Chilcotin Mountain Range, which forms part of British Columbia’s Coast Mountain Range. The landscape is characterized by the low rounded summits of the Chilcotin Range and moderately sloping upland. The property is located within the Fish Creek and Fish Lake watershed in a broad valley with slopes of moderate relief. Elevations at the site range between 1,450 m and 1,600 m above sea level. Fish Creek, in the valley bottom, flows into the Taseko River at the north end of the property. The local drainage pattern is typically dendritic.

5.4 Climate

Regionally, the Chilcotin Mountains record lower temperatures and more precipitation than the Fraser Plateau. The warmest and driest lands are generally in the main river valleys.

The Fish Lake area has a moderate continental climate with cold winters and warm summers. Local climatic conditions are moderated primarily by elevation, aspect, physiography, and the proximity of the area to the Chilcotin Mountains.

Climate estimates for the Prosperity project site have been determined using long-term regional data sources and short-term site-specific records.

The low and high estimates of mean annual precipitation are 380 mm and 527 mm respectively. The mean annual lake evaporation is estimated at 452 mm.

The long-term average annual temperature for the project site is estimated to be 2.0°C, with average monthly temperatures ranging from a high of 13.0°C in July and August to -10.0°C in December and January.

6. History

The Prosperity deposit was originally discovered in the early 1930’s by prospectors E. Calep and C. Vick, who conducted trenching of feldspar porphyritic dykes with stringers containing copper and gold values about 1.5 km east of the centre of the porphyry deposit as it is now known. In the late 1950’s, George Renner did additional work on gold-silver-copper mineralized shear zones located northeast of the deposit. In 1960 Phelps Dodge Corporation located float and subcropping mineralization that indicated a porphyry environment. The company subsequently carried out a program of induced polarization (IP), geochemical and magnetic surveys, and hand trenching. In 1963-64 they conducted a small diamond drilling program comprising 8 short holes north of the presently known deposit. The results were not encouraging and the mineral claims in the area were allowed to lapse.

In 1969, Taseko Mines Limited acquired the property and drilled 12 percussion holes totaling 1,265 m and 6 diamond drill holes totaling 1,036 m immediately to the south of the area where Phelps Dodge had explored. Taseko discovered significant tonnage grading 0.25% to 0.30% copper.

In 1970, Nittetsu Mining Company optioned the property from Taseko Mines Limited and completed 236 m of core drilling in 4 holes before returning the property to Taseko. In 1972, Taseko tested the property with 2 additional diamond drill holes totaling 156 m.

Quintana Minerals Corporation optioned the property from Taseko in 1973 and completed a 23-hole diamond drill program totaling 4,705 m during 1973-74. Vertical drill hole Q73-10, collared in the center of the deposit, intersected 415 m of disseminated and stockwork copper-gold mineralization at an average grade of 0.31% Cu and 0.54 g Au/t. The drill hole was completed, at a depth of 438 m, in mineralization of similar grade.

Bethlehem Copper Corp. optioned the property in 1979 and by 1981 had completed 3,225 m of percussion drilling in 36 holes and 10,445 m of diamond drilling in 37 holes.

Following the corporate merger of Bethlehem Copper Corp. and Cominco Ltd., Cominco acquired the Bethlehem option agreement on the property. Cominco continued to drill the property, completing 1,620 m of percussion drilling in 19 holes and 3,707 m of diamond drilling in 29 holes over the period 1982 to 1989.

A summary of this historical drilling is shown in Table 6-1.

		Percussion Drilling		Diamond Drilling		All Drilling
Year	Company	No. of Holes	(m)	No. of Holes	(m)	No. of Holes	(m)
1963	Phelps-Dodge	0	0	6	611	6	611
1964	Phelps-Dodge	0	0	2	112	2	112
1969	Taseko	12	1,265	6	1,036	18	2,301
1970	Nittetsu	0	0	4	236	4	236
1972	Taseko	0	0	2	156	2	156
1973	Quintana	0	0	14	2,972	14	2,972
1974	Quintana	0	0	9	1,733	9	1,733
1979	Bethlehem	14	1,106	0	0	14	1,106
1980	Bethlehem	22	2,119	0	0	22	2,119
1981	Bethlehem	0	0	37	10,446	37	10,446
1982	Cominco	19	1,620	12	707	31	2,327
1984	Cominco	0	0	5	1,003	5	1,003
1989	Cominco	0	0	12	1,997	12	1,997
Total Drilling		67	6,110	109	21,009	176	27,119

Cominco work programs also included 50 line km of induced polarization, magnetic and soil geochemical surveys. The induced polarization survey outlined a 2 km by 3 km east-west trending zone of high chargeability. Also undertaken was a limited metallurgical testwork program which focused on achieving high copper recovery, with little emphasis on gold recovery, using a conventional copper flotation.

In 1990, Cominco Ltd. reported a drill—indicated mineral resource of 208 million tonnes at an average grade of 0.23% Cu and 0.41 gpt Au to 360 m below surface. Many of the drill holes used to estimate this resource bottomed in resource grade gold-copper mineralization.

By agreement dated August 10, 1979, the Prosperity project was optioned by Taseko to Bethlehem Copper Corporation. Cominco Ltd. acquired the Bethlehem option agreement on the property with the merger of Bethlehem and Cominco. Under that agreement, Cominco was granted an exclusive option to acquire an 80% interest in the Prosperity project by giving notice to Taseko before November 30, 1984, of Cominco’s intention to proceed with commercial production from the Prosperity project. Cominco was entitled to extend its option on a yearly basis if Cominco concluded that it was not economically feasible to place the project in commercial production and if an independent consultant supported this conclusion. Cominco extended the option in 1984 and again in 1985, based on an evaluation of the Prosperity project prepared by Cominco in 1984. Cominco’s extension of the option was supported by a June 1986 report from Wright Engineers Limited of Vancouver, British Columbia. That report, based on

data obtained from mining and metallurgical studies provided by Cominco, confirmed Cominco’s evaluation that the Prosperity project was not commercially feasible at that time.

Taseko subsequently sued Cominco, arguing that Cominco had not complied with all of the terms necessary to enable it to extend the option, and specifically had not had a proper feasibility study prepared to determine the economic viability of the Prosperity project. Cominco successfully defended its position at the trial and appeal courts. Taseko and Cominco resolved their dispute by entering into a settlement agreement dated April 25, 1991 (the “First Settlement Agreement”). Cominco entered into the First Settlement Agreement in consideration of the issuance by Taseko of 1,000,000 common shares (issued over the period May 31, 1991 to March 31, 1992), and for the grant of a general release of Cominco by Taseko from the litigation claims made by Taseko against Cominco. The First Settlement Agreement provided that Taseko had a five-year option to sell the Prosperity Project, either directly or by way of a take-over of Taseko, in which event the proceeds would be split in a certain ratio with a maximum of CDN$48 million to Cominco.

By agreement dated December 1, 1993. Taseko acquired the exclusive right to purchase from Cominco all of Cominco’s residual interest in the project. Taseko acquired the balance of a 100% interest in the Prosperity Project by paying to Cominco CDN$2,000,140 from working capital and issuing to Cominco 1,636,364 common shares from treasury. Cominco sold 1,607,400 of these shares to net CDN$23 million and 28,964 shares were returned to treasury in April, 1994. As a result of the Second Settlement Agreement, Taseko acquired 100% of the Prosperity project free whatsoever of any royalties or third party interests.

There has been no production from the property.

7. Geological Setting

The Prosperity project is located within the western-most portion of the Intermontaine Belt at the boundary between the Intermontaine and Coast morphologic belts. The surrounding area is underlain by poorly exposed, Late Paleozoic to Cretaceous litho tectonic assemblages which have been intruded by plutons of Mid-Cretaceous to Early Tertiary age. The main Coast Plutonic Complex is 50 km southwest of the project area (Figure 7 -1).

The Yalakom Fault is the major fault in the region and lies to the southwest of the Prosperity deposit. Estimates of Eocene dextral strike-slip offsets for the Yalakom Fault have been postulated variously as ranging from 80 to 190 km (Tipper, 1969), 125 to 175 km (Kleinspehn, 1985) or 115 km (Riddell et al., 1993). It may have imparted some related structural controls that are important to the localization of mineralization at the deposit.

Northeast of the Yalakom Fault, feldspathic lithic sandstones, conglomerates and shales comprise most of the exposed rocks. These sedimentary rocks were correlated with the Lower Cretaceous Jackass Mountain Group by Riddell et al. (1993) and Schiarizza et al. (1993). The poorly exposed andesitic volcaniclastic and volcanic rocks that host the Prosperity deposit may correlate with a succession of andesites, tuffaceous sandstones, argillites and siltstones that crop out near the mouth of Fish Creek. Here, fossils collected from shales intercalated with the volcanic rocks were assigned Hauterivian (Early Cretaceous) ages (Riddell et al., 1993). The sedimentary rocks, which have been encountered in drill holes to the south of the Prosperity deposit, are likely of similar age. Sub-horizontal Miocene plateau basalts and non-marine sedimentary rocks of the Chilcotin Group form an extensive post-mineral cover in the immediate project area.

Figure 7 -1 Regional Geology

8. Deposit Types

The Prosperity Gold-Copper deposit subcrops under a 5 to 65 m thick blanket of surficial cover at the north end of Fish Lake. Interpretation of deposit geology (Caira and Findlay, 1994 and Brommeland et al., 1998) is based on a 1963 to 1997 drill hole data base consisting of 402 diamond drill holes totaling 150,185 m and 68 percussion drill holes totaling 6,309 m as outlined in Section 10.1

The Prosperity deposit is predominantly hosted in Cretaceous andesitic volcaniclastic and volcanic rocks which are transitional to a sequence of sparsely mineralized, volcanically-derived sedimentary rocks to the south (Figure 8-1). The andesitic volcaniclastics are comprised of coarse-grained crystal tuff and ash tuff, and thinly bedded tuff with lesser lapilli tuff. The upper eastern portion of the deposit is hosted by subvolcanic units of crowded feldspar porphyritic andesite and thick feldspar and hornblende porphyritic flows (Table 8-1).

In the western portion of the deposit, the multi-phase Fish Creek Stock has intruded into a thick sequence of andesite flows which overlay volcaniclastic rocks. The steeply south-dipping, oval quartz diorite stock, which is approximately 265 m wide by 800 m long, is surrounded by an east-west trending swarm of subparallel quartz-feldspar porphyritic dikes, which also dip steeply to the south. Together the stock and dikes comprise the Late Cretaceous Fish Lake Intrusive Complex that is spatially and genetically related to the deposit. Post-mineralization porphyritic diorite occurs as narrow dikes that cross-cut all units within the deposit. They represent the final intrusive phase of the emplacement of the Fish Lake Intrusive Complex.

The deposit area is overlain by a variably thick overburden cover consisting of Wisconsinian glacial till, Miocene to Pliocene basalt flows, and Tertiary colluvium and lacustrine sediments (Figure 8—2).

Figure 8—1 Geology at Overburden-Bedrock Interface

Table 8-1

Prosperity Gold-Copper Project Geology Codes

CENOZOIC

QUATERNARY COVER
Pleistocene Glacial Till
511	TILB	Basal Till
512	CLAYU	Clay
513	SICLU	Silt/Clay Mix
514	SILTU	Silt
515	GRAVU	Gravel
TERTIARY COVER
Miocene to Pliocene Basalt Flows
520	BSLT	Basalt
Colluvium
531	FANL	Fanglomerate — Limonitic
532	FAN	Fanglomerate
Glacial Lacustrine Sediments
541	GRAV	Gravel
542	SICL	Silt/Clay Mix
543	CLAY	Clay
544	SILT	Silt

MESOZOIC

LATE CRETACEOUS FISH LAKE INTRUSIVE COMPLEX
11	PMPD	Post Mineralization Porphyritic Diorite
12	INBX	Igneous Breccia
13	FP	Feldspar Porphyry
14	QFP	Quartz Feldspar Porphyry
FISH CREEK STOCK (QD)
15	QD3	Subporphyritic to Equigranular Quartz Diorite
16	QD2	Seriate Porphyritic Quartz Diorite
17	QD1	Heterogeneous Fine Porphyritic Quartz Diorite
CRETACEOUS SEDIMENTARY ROCKS
31	SEDS	Mudstone, Siltstone, Sandstone and Conglomerate
CRETACEOUS VOLCANIC ROCKS
25	SUBV	Crowded Porphyritic Andesite
24	FLOW	Porphyritic Andesite Flow
23	BEAT	Laminated Andesite Tuff
22	DEBF	Andesite Lapilli Tuff and Debris Flow
21	MAT	Andesite Tuff (ash tuff)
21	FAXT	Andesite Tuff (mainly crystal tuff)

The deposit is oval in plan, is approximately 1500 m long, 800 m wide and extends to a maximum depth of 880 m. A central potassium silicate alteration zone is co-extensive with the gold-copper mineralization. Along the deposit’s eastern margin, a discontinuous zone of phyllic alteration is developed at the boundary between the potassium silicate alteration zone and the surrounding propylitically altered rocks. The latter extend outward from the deposit for several hundred metres. Late stage sericite-iron carbonate alteration forms irregular zones, particularly within the potassium silicate alteration zone. Argillic alteration is localized along fault zones and overprints earlier alteration assemblages.

Pyrite and chalcopyrite are the principal sulphide minerals in the deposit. They are uniformly distributed as disseminations, fracture-fillings and sub-vertical veinlets and may be accompanied by bornite and lesser molybdenite and tetrahedrite-tenantite. The latter results in somewhat elevated levels of arsenic, antimony, and mercury in some parts of the deposit. Native gold occurs as inclusions in, and along microfractures with, copper-bearing minerals and pyrite. Late-stage pyrite-base metal veins, up to several centimetres in width, are most abundant within the upper eastern portion of the deposit.

8.1 Surficial Geology

Regional glaciation occurred most recently during the Wisconsinian (15,000 to 18,000 years before present) during which time ice moved over the low lying and undulating surface of the West Fraser Plateau in a northerly and northeasterly radial dispersal pattern (Talisman, 1997). The hummocky topography resulting from this period of glaciation is typical of that produced by an ablating ice mass, and includes kames, eskers and kettles deposited on top of earlier lodgment or basal till.

During Wisconsinian glaciation, ice movement in the vicinity of Fish Lake was from south to north (Caira and Findlay, 1994). Recent alluvial activity has cut into, and deposited sediments on the older Wisconsinian sediments. In the proposed pit area, three main types of glacially-derived overburden were recognized: glacial till, glaciofluvial material, and glaciolacustrine material.

Prior to the most recent glaciation, Chilcotin Group flood basalts were deposited regionally across over 25,000 km² in the interior plateau of south central British Columbia. These flood basalts are sandwiched between the Wisconsinian sediments above and, in the immediate vicinity of the Prosperity deposit, underlying colluvial and lacustrine sediments.

In general, east of Fish Creek and north of Fish Lake the overburden consists predominantly of a patchy and variably thick sequence (less than 10 m to 65 m) of basal till (TILB) that covers colluvium (FANL, FAN) and bedrock. A prominent 750 m long esker occurs on the east side of Fish Creek and extends south to within 250 m of the outlet of Fish Lake. The west side of Fish Creek is mainly underlain by a thick sequence of basalt flows (BSLT), which can be observed in cliffs outcropping along the bank of the creek. The basal till occurs as an irregular cover up to 22 m thick over the basalt flows which in turn are in direct contact with bedrock or overlie a variably extensive and irregularly thick (8 to 70 m) layer of colluvium (FANL and FAN). Lake sediments (SILT) occur extensively in the southern portion of the deposit adjacent to Fish Lake.

Detailed geological logging of the overburden within the proposed pit indicates that there are four major types of overburden present: glacial till (TILB, CLAYU, SICLU, SILTU, and GRAVU), basalt flows (BSLT), colluvium (FANL and FAN) and glacial lacustrine sediments (GRAV, SICL, CLAY, and SILT). This overburden sequence consists of 48% basalt, 38% glacial till, 10% colluvium and 4% sediments and varies from 0 to 65 m in thickness over the deposit, but is as thick as 155 m to the south of the deposit near Fish Lake.

8.2 Volcanic and Sedimentary Rocks

Five volcanic units and one subvolcanic unit comprise the majority (78%) of the Prosperity deposit host rocks. In order of volume within the proposed pit, they are: 32% andesite crystal, ash and lapilli tuff (FAXT), 23% porphyritic andesite flow (FLOW), 21% crowded porphyritic andesite (SUBV) and 2% laminated andesite tuff (BEAT). Andesite tuffs and flows are commonly interbedded.

The volcanic rocks present in the deposit area are atypical of the surrounding area and are likely of limited regional extent. Similar volcanic rocks outcrop near the mouth of Fish Creek 3.5 km to the north and may correlate with those of the deposit.

A sparsely mineralized, volcanically-derived sedimentary unit (SEDS) occupies the upper south/southeast portion of the deposit and comprises 4% of the proposed pit. Stratigraphically, these sediments are postulated to represent a facies change in the volcanic assemblage that outcrops near the mouth of Fish Creek.

8.3 Fish Lake Intrusive Complex

The Prosperity deposit is spatially and genetically related to the Fish Lake Intrusive Complex, which is comprised of the Fish Creek Stock, quartz feldspar and lesser feldspar porphyry dikes and post-mineralization porphyritic diorite dikes.

The Fish Creek Stock is a hypabyssal lenticular east-west trending, steeply south-dipping body of porphyritic quartz diorite (QD) that has intruded a thick sequence of volcanic rocks. It is composed of three phases, the heterogeneous fine porphyritic quartz diorite, seriate porphyritic quartz diorite and subporphyritic to equigranular quartz diorite units, that together comprise 11% of the deposit’s volume. These units are very similar in chemical composition, but differ in textural characteristics. The latter are commonly gradational; heterogeneous fine porphyritic quartz diorite can grade into seriate porphyritic quartz diorite and seriate porphyritic quartz diorite can grade into subporphyritic to equigranular quartz diorite over distances of several metres to tens of metres. The heterogeneous fine porphyritic quartz diorite and seriate porphyritic quartz diorite units also occur independently.

Quartz feldspar porphyry and feldspar porphyry dikes occur as an east-west trending, steeply south-dipping swarm centered east of the Fish Creek Stock. They comprise 7% of the deposit. The quartz feldspar porphyry units cross-cut all of the volcanic and sedimentary rocks identified in the deposit. The contemporaneity of the quartz feldspar porphyry dikes and the Fish Creek Stock is suggested by the occurrence of some units of transitional lithology, close to the border of the stock.

The entire suite of rocks (intrusive, volcanic and sedimentary) hosting the deposit is cross-cut by a series of barren, post-mineralization porphyritic diorite dikes (PMPD). The post mineralization porphyritic diorite unit comprises less than 1% of the deposit rocks.

Spatial distribution of the Prosperity deposit geological units is made in reference to the 1997 geology block model. The model is constructed over elevations of 547.5 to 1567.5 m above sea level, with level plans at 15 m vertical intervals. Typical plan views and vertical sections are shown in Figures 8-3 through 8-7.

Figure 8-3 Geology Plan View 1400m Elevation

Figure 8-4 Geology Plan View 1200m Elevation

Figure 8-5 Geology Plan View 1000m

Elevation

Figure 8-6 Geology Vertical Section 10100E

Figure 8-7 Geology Vertical Section 10100N

8.4 Alteration

Five main alteration styles have been identified at the Prosperity deposit: potassium silicate, propylitic, sericite-iron carbonate, phyllic and argillic. Alteration styles do not occur singularly in discrete zones; they commonly overlap and/or overprint each other. However, one alteration style will typically dominate over others in a given area, hence the naming of a zone specific to the dominant alteration style.

Potassium silicate alteration is the most widespread alteration within the deposit area. It forms a central east-west trending ovoid zone, which is intimately related to significant gold-copper mineralization (>0.20 gpt Au and >0.20% Cu). The zone of potassium silicate alteration is surrounded by propylitically altered rocks that extend outward for several hundred metres. Along the eastern margin of the deposit a discontinuous belt of phyllic alteration is developed in proximity to the transition between the potassium silicate and propylitically altered rocks. Late stage sericite-iron carbonate alteration forms irregular zones, particularly within the central zone of potassium silicate alteration. Argillic alteration is localized along fault zones and overprints earlier alteration assemblages.

The sequence of alteration events at the Prosperity deposit commenced with the emplacement of the Fish Lake Intrusive Complex and the development of a hydrothermal cell which caused the contemporaneous infusion of potassium silicate and propylitic alteration in zones concentric about the intrusive complex. This was followed by an episode of phyllic alteration which occurred at higher levels in the system and was the result of a mixing between fluids of the hydrothermal cell and meteoric waters. Phyllic alteration overprinted both potassium silicate and propylitic alteration in certain areas. Sericite-iron carbonate and argillic alteration, the latest events in the alteration history, were the result of the migration of late stage hydrothermal fluids and meteoric waters along structural features, resulting in the formation of secondary mineral assemblages in the host rocks which overprint all other alteration styles.

8.5 Structure

Numerous faults were intersected in drill core throughout the deposit area. Faults are usually indicated by strongly broken core, gouge, sheared textures, cataclastic textures and rarely mylonitic textures. All of the aforementioned features can occur across intervals of less than 1 cm to over 20 m. Utilizing all available data, two major faults (the QD and East Faults) have been delineated.

The QD and East Faults are subparallel, strike north-south and dip steeply to the west, becoming near vertical down-dip. They cut the central portion of the deposit and are approximately 230 m apart near surface and 330 m apart at depth. The western most of the two major faults, the QD Fault, trends approximately 355° and has a steep westward dip of 82° to 86°. This fault marks the eastern boundary of the Fish Creek Stock. The eastern most of the two major faults, East Fault, strikes approximately 360° and has a steep westward dip of 85° to 87°.

9. Mineralization

Gold-copper mineralization within the Prosperity deposit is intimately related to potassium silicate alteration and a later, superimposed sericite-iron carbonate alteration. This is particularly true within a central, east-west trending ovoid zone that hosts the majority of the mineralization.

Chalcopyrite-pyrite mineralization and associated copper and gold concentrations are distributed relatively evenly throughout the host volcanic and intrusive units in the deposit. A sedimentary unit, which is located in the upper southeastern part of the mineralized zone, is sparsely mineralized. Post mineralization porphyritic dikes are essentially barren.

Pyrite and chalcopyrite are the principal sulphide minerals and are accompanied by: minor amounts of bornite and molybdenite; sparse tetrahedrite-tennantite, sphalerite and galena; and rare chalcocite-digenite, covellite, pyrrhotite, arsenopyrite and marcasite. Native gold generally occurs as inclusions in, and along microfractures with, copper sulphides and pyrite. Pyrite to chalcopyrite ratios throughout most of the proposed pit area range from 0.5:1 to 1:1 and rise to 3:1 or higher around the periphery of the deposit which coincides with the propylitic, and locally the phyllic, alteration zones.

Sulphide minerals show the thoroughly dispersed mode of occurrence characteristic of porphyry copper deposits. Sulphides occur in relatively equal concentrations as disseminations, blebs and aggregates in mafic sites, as fracture fillings and as veinlets. Disseminated sulphide mineralization is marginally more prevalent than veinlets in intrusive rocks while in volcanic rocks the reverse was noted.

Gold and copper distributions throughout the deposit are presented in Figures 9–1 through 9–3.

Figure 9—1 Au and Cu Grades at 1400m Elevation

Figure 9—2 Au and Cu Grades at 1200m Elevation

Figure 9—3 Au and Cu Grades at 1000m Elevation

10. Exploration

10.1 Extent of All Relevant Exploration

Up to 1991, exploration programs at the Prosperity Project included extensive IP, magnetic and soil geochemical surveys, and 176 percussion and diamond drill holes totaling approximately 27,100 m as outlined in Section 6. This work helped define the Prosperity Project mineralization to a depth of 200 m, and outlined a gold-copper mineralized zone approximately 850 m in diameter which Cominco estimated as a geological resource of 208 million tonnes grading 0.23% copper and 0.41 gpt Au.

In 1991 Taseko drilled 10 holes totaling 7,506 m in a “cross” pattern to test the core of the deposit over a north-south distance of 550 m. All of the holes intersected continuous significant copper and gold grades and extended the mineralization to 810m below surface. A scoping-level metallurgical test work program was completed by Melis Engineering Ltd. The test work demonstrated that acceptable gold and copper recoveries could be achieved by bulk sulphide flotation followed by regrinding and conventional copper flotation. Baseline environmental and monitoring studies were initiated by the Company.

Diamond drilling continued in 1992, and by the end of the year an additional 116 HQ and NQ diameter vertical drill holes totaling 60,558 m had been drilled, expanding the deposit to 1400 m east-west, 600 m north-south and to 850 m below surface. G. Giroux, P.Eng., reported mineralized material (unclassified mineral resource) of 976 million tonnes at an average grade of 0.23% Cu and 0.48 gpt Au.

In 1993, eight holes totaling 2,100 m were drilled for geotechnical purposes.

Subsequent to 1993 comprehensive metallurgical tests by Melis Engineering Ltd. and a 1994 pre-feasibility report by Kilborn Engineering Pacific Ltd., the Company completed a 12 hole (4,605 m) inclined core drilling program in 1994 to investigate the distribution of fracture controlled gold and copper mineralization in the deposit. In addition, 22 holes (3,171 m) were drilled to investigate geotechnical conditions in the proposed Project development areas.

In 1996 and 1997, an additional 107 holes (49,465 m) were completed in order to upgrade the confidence limits of the deposit. Of this total, 20 holes (2,203 m) were drilled vertically and 87 holes (47,262 m) were inclined. These holes significantly increased the density of pierce points in the deposit and added to the geotechnical and geochemical characterization of the rock in the deposit.

Over the 34-year period from 1963 to 1997, a total of 154,631 m has been drilled in 452 holes on the Prosperity project. Of this total, 210 holes (77,173 m) were core drilled vertically and 174 holes (71,149 m) were inclined. Sizes of cored holes have included BQ, HQ, and NQ totaling 148,322 m, with an average drill spacing of 70 m. The balance of 6,309 m is from 68 vertical percussion drilling holes. A summary of the drilling of this period is shown in Table 10-1 and Figures 10-1 and 10-2.

Table 10-1

Drilling Summary 1963 – 2007

		Percussion Drilling		Diamond Drilling		All drilling
Year	Company	No. of Holes (m)		No. of Holes (m)		No. of Holes (m)
Pre 1990	Section 6.1	67	6,109	109	21,009	176	27,118
1991	Taseko	0	0	10	7,506	10	7,506
1992	Taseko	0	0	116	60,558	116	60,558
1993	Taseko	0	0	8	2,104	8	2,104
1994	Taseko	1	200	34	7,680	35	7,880
1996	Taseko	0	0	69	28,423	69	28,423
1997	Taseko	0	0	38	21,042	38	21,042
1998*	Taseko	0	0	18	1,768	18	1,768
2007*	Taseko	5	44	6	1,762	11	1,806
Total Drilling		73	6,353	408	151,852	481	158,205

In 1998, G. Giroux, P.Eng., reported estimated measured and indicated mineral resources of 1.0 billion tonnes at 0.41 gpt Au and 0.24% Cu and an inferred resource of 0.2 billion tonnes grading 0.25 gpt Au and 0.21% Cu at a 0.14% copper cut-off. Giroux provided the resource at a number of different cut-offs, and since that time, the resource has also been reported at a 0.2% copper cut-off, which was also based on the Giroux 1998 estimate.

* 1998 drilling consisted of fourteen geotechnical holes and four in-pit verification holes. The 2007 drilling included six in-pit metallurgical holes and five geotechnical holes outside the pit. These holes were not incorporated into the Geological model.

Figure 10-1 Drill Hole Locations 1969 to 1994

Figure 10-2 Drill Hole Locations 1996 to 1998

11. Drilling

11.1 Drilling Pre-1991

Prior to 1991, several companies unrelated to Taseko carried out mineral exploration in the Prosperity project area. The first drilling on the property was carried out by Phelps Dodge in 1963 and 1964. Taseko Mines (Old Taseko), Nittetsu Mining Company, Quintana Minerals Corporation, Bethlehem Copper and Cominco also drilled holes between 1969 and 1991. During the 13 years of active exploration during this period, approximately 21,009 m of core drilling in 109 holes, and 6,109 m of percussion drilling in 67 holes was completed.

11.2 Drilling 1991-1994

In 1991, the current Taseko management, (New Taseko), took over the project and from 1991 through 1994 a total of 77,848 m was cored in 168 holes. Approximately 28% of the drilling was HQ (6.35 cm diameter) core size and 68% NQ (4.76 cm diameter) core. A single percussion hole 200 m in length was also drilled.

The bulk of the drilling took place in 1991 and 1992 as a series of predominantly vertical drill holes. In 1993 four drill holes inclined at –60° were drilled using oriented core methods and four vertical holes were completed. The single percussion hole and 34 core drill holes were drilled in 1994. The latter included eight holes at various orientations, drilled outwards from the centre of the deposit. Another 13 holes were drilled in the main deposit area at an azimuth of 340° with inclinations of -45° to -50 °. A further 13 holes were drilled vertically in 1994. No drilling took place in 1995.

Geotechnical data was recorded for all drill holes from 1991 to 1994. Core recovery was measured on 25,344 drill run intervals averaging 3.05 m in length. Recovery was good, with a mean value of 97.0% and a median value of 99%, for the sampled intervals measured.

11.3 Drilling 1996-1997

From June 1996 through May 1997, Taseko completed a 107 hole drill program, comprising 49,465 m of diamond bit core drilling. This included in-fill definition holes, oriented in-fill holes, oriented geotechnical pit wall holes, acid base accounting (ABA) holes, and waste rock and tailings geotechnical holes (Table 11-1). All in-fill drilling was performed at an azimuth of 340° East and an inclination of -45°, on 100 m by 100 m spacing. Of this drilling, all was sampled and assayed except 4,400 m of overburden.

JT Thomas Diamond Drilling of Smithers, BC drilled HQ and NQ core by using skid-mounted hydraulic drills and a drill modified for helicopter transport to remote sites. Geotechnical data was recorded for all but six drill holes in the 1996-1997 programs. Core recovery was measured on 17,035 drill run intervals averaging 3.05 m in length. Recovery was good, with a mean value of 94.2% and a median value of 98% for the sampled intervals measured.

Table 11-1

Drill Holes by Orientation by Year

		1963 to	1991,			1996,
Drill Type	Orientation	1990	1992	1993	1994	1997	1998	2007	All

Percussion	Vertical	67	—	—	1	—	—	5	73
Core	Vertical	58	115	4	13	20	15	—	225
Total	Vertical	125	115	4	14	20	15	5	298
	North	25	2	—	1	—	—	—	28
	Northeast	1	1	—	2	1	—	2	7
	East	15	6	2	—	—	—	—	23
	Southeast	2	—	—	1	1	—	—	4
Core	South	1	—	—	1	5	—	—	7
	Southwest	3	—	1	1	—	1	1	7
	West	—	2	—	1	1	—	1	5
	Northwest	4	—	1	1	—	—	—	6
	Azim. 340°	—	—	—	13	79	2	2	96
Total	Inclined	51	11	4	21	87	3	6	183
Total	All	176	126	8	35	107	18	11	481

11.4 Drilling 1998

Eighteen core holes were drilled in 1998 including 15 vertical holes and three holes inclined at -45° for a total of 1,768 m. Four drill holes were completed within the main porphyry, nine geotechnical holes were drilled to the south of Fish Lake, and five geotechnical holes were drilled east of the main deposit area.

In 1998 Kilborn undertook a comprehensive audit and verification program of the geology and assay results of the Prosperity project. The drilling noted above was in support of this program. The four in-pit verification diamond drill holes totaling 1,150 metres were completed by Kilborn. 110 half core samples and 99 reject samples from the 1991-1992 drill programs were re-assayed. All analytical work was performed by Chemex. Based on the results of this program, it was Kilborn’s opinion that the geological work for the Prosperity was done in a professional manner and according to industry standards.

11.5 Drilling 2007

Six core holes totalling 1,762 metres were drilled in 2007 for metallurgical purposes. The purpose of this program was to provide sufficient material representative of the various alteration zones anticipated to be encountered in the first six years of production. Based on detailed comparison of the geological information predicted by the pre-2007 geological model and the actual intersections in these drill holes, there is a high degree of confidence in the existing model. Limited local variance in the alteration or lithology between the predicted and observed is attributable to portions of drill holes passing through the very edges of the modeled units.

Overall, there is very good agreement between the drilled and the predicted intersections, lending credence to the existing model.

12. Sampling Method and Approach

The Prosperity deposit was explored and extensively drilled by seven different companies between 1963 and 2007. A total 158,204 m of core and percussion drilling was completed in 481 drill holes during the twenty one years in which active drill exploration took place (refer to Sections 6 and 10). The drill hole spacing is such that no part of the deposit, as defined by the current resource model, is farther than 70 m from drill hole information, although the majority of drill holes are considerably closer.

A total of 65,489 drill core samples and 1,674 percussion samples were taken for analysis between 1969 and 2007. Sampled and assayed intervals total 141,087 m. For most holes drilled, the entire length of Cretaceous rock was sampled and assayed. Early sampling and analysis at Prosperity focused on assessing the copper mineralization visible in the rocks. Once the presence of significant gold mineralization was recognized, assaying for gold became more comprehensive. Starting in 1991, Taseko undertook multi-element analysis for 30 elements on all drill core samples, in addition to the regular assaying for copper and gold.

12.1 Core Logging

All drill holes completed from 1991 through 1998 in the main deposit area were geotechnically logged, geologically logged, and photographed prior to sampling.

12.2 Sampling

A total of 64,511 drill core samples and 1,548 percussion samples have been taken for analysis since 1969. Prior to 1991, a total of 6,905 were taken with an average length of 3 m. During the 1991 through 1998 drilling programs, 58,580 core samples were taken for assay. These sample intervals were generally 2 m in length, except in instances where this was impractical. No assay information exists for the eight holes Phelps Dodge drilled in 1963-1964. Table 12-1 lists the number samples and type of analysis by year.

During the period 1991-1994, drill core was mechanically split, one half of which was submitted for preparation and analysis. Of the total meterage drilled during 1996-97, 42% was subject to whole core sampling, 44% was sampled as sawn half-core, 5% of samples comprised the larger portion of core sawn 80:20. The remaining 9% was cored overburden, which was not generally sampled, although some samples were taken for ABA studies and placer claim assessment. In 1998 the samples were half sawn core. For the 2007 metallurgical program, the core was sawn in the proportion 80/20 for metallurgy/assay respectively. Samples were sawn lengthwise with a diamond bladed rock saw using water to wash and lubricate the blade. The remaining sample was put back in drilling order in the core box. Figure 12-1 illustrates the sampling, sample preparation and analytical protocol for the 1996-1997 programs. Drill core remaining after sampling was returned to the core boxes, which were racked and stored at the Prosperity Site.

Table 12-1

Number of Samples by Year

	Au & Cu	Au Assays	Cu Assays	All
Year	Assays	Only	Only	Assays
1969	120	—	542	662
1970	22	2	22	46
1972	37	—	—	37
1973	706	6	3	715
1974	486	—	—	486
1979	76	32	119	227
1980	442	37	90	569
1981	2,333	61	347	2,741
1982	452	53	83	588
1984	267	—	—	267
1989	564	3	—	567
1991	3,472	—	—	3,472
1992	28,700	—	—	28,700
1993	581	—	—	581
1994	2,744	—	—	2,744
1996	12,724	—	—	12,724
1997	9,606	—	—	9,606
1998	539	—	26	565
2007	761	—	—	761
Total	64,633	194	1232	66,059

Figure 12-1

1996-1997 Drill Core Sampling, Preparation and Analytical Flow Chart

13. Sample Preparation, Analyses and Security

13.1 Security

In 1991-2007 the drill core was boxed at the drill rig and transported twice daily by company truck to the logging, sampling and sample preparation compound at the Prosperity site. The core was geologically and geotechnically logged, given quality control quality assurance (QAQC) designations, photographed and sampled under the supervision of Taseko geological and engineering staff. Samples were placed in shipping sacks and taken by company truck to Williams Lake and then shipped by commercial carriers to the Vancouver area analytical laboratories.

13.2 Sample Preparation

During the 1991-2007 programs, core samples were shipped to Vancouver laboratories for preparation, including drying at temperatures less than 65°C, and blind standards inserted by the preparation laboratory into the sample stream, crushing and pulverizing. Beginning in 1994, all samples were weighed to the nearest 10 grams and blind standards were inserted into the sample stream by the preparation laboratory. An average dry weight of 7.4 kg was obtained from the 24,804 weights reported for the 1994-1997 samples. The 761 samples from the 2007 program had an average dry weight of 3.4 kg. The 1991-1994 samples were prepared by Mineral Environments Laboratories Ltd. (Min-En) of North Vancouver BC. In 1996-1997, the samples were prepared by either Acme Analytical Laboratories Ltd. (Acme) or CDN Resource Laboratories Ltd. (CDN) of Vancouver BC. The 1998 and 2007 samples were prepared by ALS Chemex of North Vancouver BC.

Primary comminution to approximately ¼ inch (6.4 mm) size was provided by a jaw crusher in 1991-1993. A secondary roll crusher was used to obtain minus 15 mesh (0.125 mm) material for pulverization. In 1994-2007 the dried samples were crushed in a single stage so that more than 60-70% passed a 10 mesh (2mm) screen. Since 1991, a sub sample (the assay split), weighing a minimum of 500 grams, was riffled from the crushed material. The remaining crushed reject material from the 2007 program is stored in Taseko’s Surrey BC warehouse All previous coarse reject material was discarded in 2001.

Preparation of the assay splits involved ring and puck pulverization. The 1991-1993 laboratory specifications for pulverization were approximately 95% passing 120 mesh (0.125 mm). In 1994-1997 the specifications for pulverization were modified to greater than 90% passing 150 mesh (0.1 mm). In 2007, the samples were pulverized to 85% passing 75 micron. Screen tests were done and reported for approximately one in fifty pulps. Additional, detailed screen analyses were done on selected samples by CDN and International Metallurgical and Environment Inc. After testing, plus and minus fractions of the screen samples were recombined, and the samples kept within the normal sample stream.

In 1991-1993 the 500 g pulp was homogenized by rolling prior to analytical aliquot selection. Starting in 1994, and continuing through the 1996-1997 programs, sample preparation was carried out separately from the assaying and analytical work, and reported on separate laboratory certificates. In these years, each 500 g pulp was rolled prior to riffle splitting and a 125 g analytical sub sample was obtained. This was placed in a pulp bag bearing the sample number

and lot code and shipped to the analytical lab. In 1998 and 2007, sample preparation and analysis was performed at the same laboratory on the same workorder. Pulps remaining after splitting and aliquot selection were returned to Taseko and stored in the company warehouse.

13.3 Sample Analysis 1991-2007

Min-En performed the primary analytical work from 1991 through 1997. Gold was analyzed by lead collection Fire Assay, using a one assay ton (30 g) charge. After fusion, the doré bead was finished by Atomic Absorption Spectroscopy (AAS). Gold assays greater than 10 gpt were automatically re-assayed by one assay ton Fire Assay fusion with a gravimetric finish. Copper was determined by Aqua Regia digestion on a 0.5-2.0 gram sample with an AAS finish. In addition, 0.5 gram aliquots of all samples were assayed for 31 elements by Aqua Regia digestion Inductively-Coupled Plasma Atomic Emission Spectroscopy (ICP-AES). Mercury determinations were performed by Cold Vapor AAS.

ALS Chemex performed the primary analytical work on the samples from the 1998 program submitted directly by Kilborn Engineering. The 1998 samples were analyzed by the same methods as the 1991-1997 samples. In 2007, the samples were analyzed by ALS Chemex for: Au, Pt and Pd by 30 g Fire Assay ICP-AES finish, Cu by four acid digestion with an AAS finish, Hg by cold vapour AAS finish, 48 elements by four acid digestion ICP-MS finish, and whole rock analysis by lithium borate fusion XRF finish.

The following conversion factors were used for the analytical results:

· 1 gpt = 1,000 ppb

· 1 % = 10,000 ppm

· 1 oz/Ton = 34.2857 gpt

13.4 Quality Assurance Quality Control

Taseko Mines Limited implemented a quality control quality assurance (QAQC) program after taking over the Prosperity project in 1991. This program was in addition to the QAQC procedures used internally by the analytical laboratories. The results of this program indicate that analytical results are of high quality and suitable for use in detailed modeling and resource evaluation studies. Table 13-1 describes the QAQC sample types used in this program.

Table 13-1

QAQC Sample Types Used

QC Code	Sample Type	Description	Percent of Total
MS	Regular Mainstream	· Regular samples submitted for preparation and analysis at the primary laboratory.	90%

ST	Standard Reference Material	· Mineralized material in pulverized form with a known concentration and distribution of element(s) of interest · Randomly inserted using pre-numbered sample tags	5% or 1 in 20

DP	Duplicate or Replicate	· An additional split taken from the remaining pulp reject or coarse reject. · Random selection using pre-numbered sample tags · Inter-Laboratory duplicates analyzed at a second or check laboratory (random selection) · Non-random selection, after initial assays returned	10% 1 in 10 (1991- 1992) 5% 1 in 20 (1996- 2007)

SD	Standard Duplicate	· Standard reference sample submitted with duplicates and replicates to the check laboratory	<1 %

Table 13-2 is a summary of the regular mainstream (MS) samples and additional QAQC samples analyzed on the Prosperity Project that were submitted by Taseko in addition to the laboratory internal QAQC work. Only a limited number of QAQC samples exist prior to 1991.

Table 13-2

Drill Hole Sample QAQC Summary

Year	MS	DP	SD	ST	Total
Pre-1991	6,905	109	—	—	7,014
1991	3,472	351	—	—	3,823
1992	28,700	2,819	—	—	31,519
1993	581	—	—	—	581
1994	2,744	73	11	131	2,959
1996	12,724	677	46	636	14,083
1997	9,607	499	33	467	10,606
1998	565	284	5	25	879
2007	761	41	4	34	843
ALL	66,059	4,853	99	1,293	72,304

Standards

In 1994, Taseko modified the sampling and analytical QAQC program, to include the random submission of project-based, bulk standard reference materials within the mainstream and duplicate analytical streams. The insertion of standards continued over the course of the 1996-2007 drill programs. The property standards were inserted by the sample preparation laboratory approximately mid-way between duplicate samples. On average, one in twenty samples was randomly selected for duplicate analysis at a second laboratory. In addition, approximately every thirteenth standard was designated as a standard duplicate, to be submitted with the duplicate sample stream at the second laboratory.

This process involved identifying the QAQC samples at the core logging stage. Sample bags containing ‘standard’ tags (but no sample) were inserted at the appropriate intervals at this stage. Standard tags were numbered as part of the normal sample sequence. Quality control samples were also identified on the sample shipment notice and marked on the bags in the same fashion as regular samples.

The 1996-1998 program employed four property-based standards: 94FLH1, 94FLM1, FL96M2 and FL96L1, with designations H, M and L corresponding to high, medium and low gold grades respectively. Standard preparation took place at CDN under the direction of Smee & Associates Consulting Ltd. (Smee) The material was crushed, pulverized, and screened to minus 150 (106 microns) in 1994 or minus 200 mesh (75 microns) in 1996 and then mechanically mixed. As mixing proceeded, sub-samples were periodically assayed to test for homogeneity. Depending on the standard, test results indicate that this was achieved within 2 to 5 days. In 1994 the assessment for homogeneity took place in-house, whereas in 1996, the sub-sample results were forwarded to Smee for verification.

In 2007, commercial copper-gold certified reference materials (CRM) were employed as assay standards.

For each standard, the sample preparation laboratory received an otherwise empty bag bearing the tag or number of the designated sample and the standard reference code. Prior to shipping the 125 g pulps prepared from the regular samples, 125 g standard pulps were inserted into the sample stream. These had the same appearance as the mainstream samples and fitted sequentially into the sample number series in order to appear anonymous to the environmental laboratories.

Gold and copper analytical results were monitored when received for QAQC failures, including results outside the control limits, or consecutive results outside the warning limits.

· Warning limits: ± 2 S.D.

· Control Limits: ± 3 S.D.

Table 13-3 lists the four project based standard reference samples used in the 1994-1998 programs and the two commercial CRM used in 2007.

Table 13-3

Summary of Copper-Gold Standard Reference Materials Used

	Times		2 Std.		2 Std.
Standard	Used	Cu %	Dev. Cu	Au gpt	Dev. Au
94FLH1	255	0.381	0.038	0.742	0.120
94FLM1	337	0.253	0.054	0.395	0.099
FL96M2	150	0.269	0.023	0.372	0.077
FL96L1	501	0.166	0.010	0.259	0.034
CGS-12	24	0.265	0.015	0.290	0.040
CGS-16	7	0.112	0.005	0.140	0.046

Most of the larger spikes in the standard performance charts are attributable to the mislabeling or accidental insertion of a different standard by the sample preparation lab. If an assay of a standard returned outside of the set tolerance limits, the pulps of the mainstream samples bracketing the standard were re-assayed at Min-En. As a result, 438 pulp samples were re-run, the results of which are in very close agreement with the original assays. Depending on the behavior of the standard in these check batches versus the original runs, the original results were replaced in the database by the second result. In summary, the assay performance based on the standard results is within acceptable limits.

Blanks

A total of 19 granite blanks were inserted in 2007 to monitor for contamination during sample preparation and analysis. No contamination was detected. Blanks were not inserted with the samples submitted by Taseko prior to 2007. Analysis of drill core made up entirely of material from the Post Mineral Dyke unit, which commonly occurs within mineralized porphyry rocks

consistently returned values <0.05 g/t gold and <0.05% copper. Analytical blanks were regularly inserted by the assay laboratories and analyzed with each batch of samples as part of their internal QAQC protocols.

Duplicates 1991-1992

Prior to 1991, a total of 82 duplicate gold assays and 27 duplicate copper assays were completed, representing 1.4% of the total gold assays and 0.4% of the total copper assays. A more thorough check assay program of random duplicate analysis was implemented starting in 1991-1992. Every tenth two metre sample from the 121 holes drilled during this period, regardless of grade, was shipped to Chemex for riffle splitting of the coarse reject, pulverization, and analysis for gold and copper. Then, for each of the duplicate samples analyzed, Chemex took another riffle split from the coarse reject, which was re-bagged, renumbered, and re-submitted to Min-En for pulverization and “blind” gold-copper analysis.

Thus 3,171 samples, representing 10% of the 1991-1992 programs, were analyzed three times for gold and copper. There is good agreement in the intra-laboratory and inter-laboratory reject duplicate results for copper and gold through all grade ranges for the 1991-1992 programs.

Duplicates 1994 and 1996-1997

Similar duplicate programs were conducted in 1994 and 1996-97, although as noted above, sample preparation was not carried out at the same laboratory as the final assays. Duplicates were therefore held at the preparation laboratory until batches of twenty or more had accumulated in order to ensure that each batch shipped to the assay laboratory was accompanied by a standard. International Plasma Laboratory Ltd. (IPL) of Vancouver performed the check assay analysis in 1996-1997.

In 1994 a total of 73 duplicate and 11 standard duplicate samples were analyzed following this procedure. During the same year, 1,841 vein samples, assayed for copper and gold, were taken from the other half of the core. Consequently fewer reject duplicates were taken relative to the 1996-1997 program when a total of 1,146 reject duplicate and 79 standard duplicate samples were assayed, following the above procedure. In addition, 4,389 vein samples were assayed for copper and gold during 1996-1997.

In 1996-1997 the sample prep laboratory (Acme or CDN) prepared a duplicate 500 g sub sample. This sample was riffled from the —10 mesh crushed reject after the mainstream sub-sample had been taken. These duplicates were then pulverized in the same manner as mainstream samples. Duplicates were held at the preparation laboratory until batches of twenty or more had accumulated before they were sent to Chemex in 1994 and to IPL in 1996-1997 for check assay analysis. This procedure ensured that the duplicates were also accompanied by a standard.

There is good agreement in the 1996-1997 inter-laboratory reject duplicate assays for copper and gold through most of the grade ranges. Some differences in the Min-En versus IPL results appear below 0.1% copper and 0.1 gpt gold, which likely reflect different analytical detection limits at the two labs.

Results of the duplicate analysis are well within the anticipated range for inter-laboratory checks.

Duplicates 1998 and 2007

Section 14.2 on verification describes the results of the independent review by Kilborn on the analysis of number of different types of duplicate samples in 1998.

In 2007, a total of 41 coarse reject duplicates were pulverized and assayed by Acme Analytical Laboratories using similar methods to ALS Chemex. Comparison of the matched pair results shows very good correlation for Cu and good correlation for Au.

13.5 Specific Gravity — Bulk Density Measurements

In 1991-1992, specific gravity measurements were carried out on-site at regularly spaced 8 metre increments on 8 to 15 cm long pieces of core by the water immersion method. Additional specific gravity measurements were taken in 1996-1997. Specific gravity determinations were made from 7,687 representative field measurements. Laboratory determinations and laboratory checks were also made. Table 13-4 is a summary of the specific gravity measurements by year and method.

The model developed by Taseko accounts for variation in specific gravity with alteration type, lithologic type, and geographic domain. A bulk density model was created by applying a reduction factor to the modeled specific gravity to account for the estimated bulk void space. This followed a recommendation in a 1993 Knight Piésold Ltd. memo Report on Influence of Geotechnical Factors on Bulk Density. Hence an upper limit reduction factor of 0.25% was applied to values above the gypsum line while a reduction factor of 0.125% was used below this level.

Table 13-4

Specific Gravity Measurements by Year and Method

					Laboratory
	Field		Laboratory		Check
Year	No.	Average	No.	Average	No.	Average
1991	811	2.73	95	2.71	11	2.70
1992	6,753	2.73	430	2.71	48	2.71
1996	25	2.71	—	—	—	—
1997	98	2.74	—	—	—	—
Total	7,687	2.73	525	2.69	59	2.71

14. Data Verification

14.1 Database

In 1991 Taseko acquired the pre-1991 drill hole data from Cominco, including digital logs and a partial set of drill hole files. In most cases the original logs and laboratory analytical certificates were not available, although some half core material and crushed rejects were still accessible. While these were generally in such poor condition as to be unusable for the purposes of re-assaying, enough material was available from some of the pre-1991 holes for Taseko to re-log them in 1991-1992. These logs confirmed the overall geological interpretation and the tenor of copper mineralization in these cores. Cominco’s digital compilation of pre-1991 information was reviewed and routine errors and omissions were corrected. For the most part, the Cominco compilation was accepted at face value.

Since 1991, all drill logs, sampling and analytical information has been compiled in an Access relational database which has tables that are compatible with GEMS mining exploration software. Prior to 2007 written logs were produced manually at the logging compound and site offices. The key drill data tables: header, survey, geotechnical, geology, vein, density and sample description were entered into spreadsheets at the project site and Vancouver office and imported into the database. In 2007 logs were entered directly into an Access database designed for this purpose. The field data was merged with the analytical results in the Vancouver office and the compiled information was then exported to Vulcan and MineSight for further processing and modeling. A complete set of core photographs was taken. The prints and negatives are archived in the project files.

The database for the current block model and resource estimate is based on drilling completed by Taseko Mines Limited from the period 1991-1997, and also includes 87 pre-1991 drill holes considered to be most reliable. This additional diamond drill data assisted with geological interpretation and provided extra control during block grade estimation.

The pre-1991 percussion drill holes were not included since the sampling lengths and procedures were quite different from diamond drill core sampling and there was no lithologic control in the sampling of percussion chips. Four holes drilled by Nittetsu in 1980 were also excluded as the distribution of both copper and gold grades for these holes were clearly different and therefore the assays were considered suspect.

A series of post-1991 holes on the edges of the deposit, drilled primarily for geotechnical information, were also incorporated in the database to define the outer boundary of the resource and the proposed pit. In these seven drill holes, yielding a total of 429 samples, copper grades were by ICP-ES analysis and gold grades were by one half assay ton (15 g) Fire Assays.

Drill hole collars at the Prosperity Project were located with reference to the Taseko Mine Grid, which was originally established for soil geochemistry and mapping purposes in the 1960’s. Bethlehem surveyed several baseline hubs and claim posts in this co-ordinate system in 1979, and in 1981 McElhanney Associates tied the Mine Grid into government triangulation stations ‘Tex’ and ‘Junior’. At this time, McElhanney also established 228 control points for photogrammetric mapping purposes, provided the first Mine Grid elevations, surveyed the

locations of 13 claim posts, and surveyed 111 drill hole collars. Mine Grid co-ordinates and elevations of other pre-1991 drill holes, which had not been surveyed, were provided by previous operators of the project.

In 1993, the pre-1991 collar locations and elevations were checked by plotting them on the McElhanney 1:2500 scale 2.5 m contour map. The plotted elevations were then compared with the mapped contour elevations to check for discrepancies in elevation. Most differences were less than +/-2 m, however 11 holes had differences of up to +/-5 m, which were not resolved.

In 1991-1992 site surveying was undertaken by Taseko based on the control provided by the 1981 McElhanney mine-grid survey stations 1086, 1087 and 1098. These control stations were verified by McGladrey Surveys Ltd. in the spring of 1993 (Caira and Findlay, 1994). McGladrey also resurveyed 42 drill hole collars, and the co-ordinate comparison between the McGladrey and Taseko surveys was very good. McGladrey noted mean differences in northing, easting and elevation of 0.0, -0.1 and 0.2 m respectively between the two sets of collar survey results.

A Total Station survey instrument was used to survey the twenty-nine 1994 drill hole collars located within the main deposit area. The other six 1994 holes, which were drilled south of Fish Lake along the proposed tailings embankment were surveyed to +/- 1m using a differential GPS in 1997.

During the course of further photogrammetric and orthophoto work in 1996, McElhanney re-checked the key survey stations and provided a method to convert from Mine Grid to UTM NAD 83 Zone 10 co-ordinates. In 1996-1997, new drill holes located within the main deposit area were surveyed using a Nikon total station and TDS 500 Data Collector. Data was downloaded directly to a computer spreadsheet in order to eliminate transfer errors. These results are summarized in the Prosperity Site 1996-1997 Survey Report (Maguire, 1997). Outside of the main deposit area, collar coordinates were obtained with a Trimble Scoutmaster GPS. UTM coordinates were transferred manually from this instrument to a spreadsheet.

Down hole surveys were taken predominantly on non-vertical cored drill holes in the deposit area between 1981 and 1992. The measurements were taken at approximately 150 metre intervals using magnetic tools. From 1993 to 1998 the Light Log down hole survey tool with an average spacing between measurements of 3 metres was used. In 2007, surveys were taken with a Reflex EZ-shot magnetic compass tool approximately every 60 metres down each cored hole.

14.2 Verification

Taseko verified the post-1990 portion of the Prosperity drill hole database by manual team verification in late 1992 and early 1998. This work focused on the following areas: sample logs, assay results, laboratory measured specific gravity measurements, collar and down hole surveys and geology. In addition to this, all drill hole copper and gold assay and geologic data was plotted out in 15 m level plans and 50 m spaced cross-sections and visually validated by the project geologists.

A report on validation of drill hole orientation information from 1991-1997 was completed by Taseko in 1998. Drill hole orientation is derived from surface surveys of the collar azimuth and dip, Light Log tool downhole readings, single shot Sperry Sun magnetic down hole compass measurements, and a few downhole acid-test readings of dip. Corrections were applied to the database in the case of miss-readings. Survey records where tool malfunction was suspected were removed.

The results of the Taseko verification program indicate that the database is of good quality and acceptable for use in geological and resource modeling

As part of the feasibility study, Kilborn undertook a comprehensive audit and verification program of the geology of the Prosperity project in 1998. This included: a survey audit, a geological audit, a check assay program, a drilling, sampling and assaying program; and an additional sampling program.

As part of this program Kilborn completed a verification drilling, sampling and analytical program with logistical support provided by Taseko. Four diamond drill holes, 98-286, 98-287, 98-288 and 98-299 totaling 1150 m were drilled by Major Drilling Ltd. Drilling was done in NQ2 except one twin hole (98-289) which was partially drilled as HQ to a depth of 153 m in order to exactly replicate the Taseko hole (92-26). All drilling and sampling work was supervised by Kilborn personnel, who also delivered the samples to the assay laboratory.

The check assaying, sampling and drilling programs were designed to increase the confidence level with respect to grades and geological interpretation utilized in resource and reserve estimation for the Prosperity project in areas to be mined mostly during the payback period. Although some variation in grades and geological descriptions were noted, it was felt that they were random and not of significance. Based on the results of this program, combined with the rest of the geological audit/verification program, it was the opinion of Kilborn that the geological work for the Prosperity deposit was done in a professional manner and according to industry standards.

15. Adjacent Properties

There is no information of significance relative to the Prosperity project from adjoining property.

16. Mineral Processing and Metallurgical Testing

16.1 Introduction

Early metallurgical test work on the Prosperity Project was conducted by various laboratories from 1973 to 1991. The early test work was limited in scope and nature, primarily being focused on achieving high copper recovery at medium to high concentrate copper grade, with little emphasis on gold recovery to the final copper concentrate. An initial Phase I metallurgical test program was undertaken by Melis in 1991 on composites made up from upper level assay reject samples taken from a 1989 drilling program. Although the composites tested were not deemed representative of the Prosperity mineralization, this initial test program did show that acceptable copper and gold recoveries could be achieved using a bulk sulphide float at natural pH with subsequent copper-pyrite separation and concentrate cleaning at alkaline pH.

A more extensive second phase of metallurgical testing was completed on representative composites of Prosperity mineralization at Lakefield Research Limited from December 1992 to August 1993 under the supervision of Melis (Melis Project No. 265; January 25, 1994 Report). This Phase II test program included batch flotation tests and locked cycle flotation tests on composites representing different areas of the deposit to determine achievable copper and gold recoveries and provide detailed concentrate analysis, grindability assessments, tailings settling tests, environmental data, and a cursory examination of the removal of mercury, arsenic, lead and antimony from Prosperity flotation concentrate.

A third phase of metallurgical testing was a run-in pilot plant carried out on Prosperity assay rejects and half (drill) core combined composites from the 1994 drilling program at Lakefield Research Limited in November, 1996 under the supervision of Melis (Melis Project No. 333; March 25, 1997 Report). The run-in pilot plant was carried out as a precursor to the later comprehensive pilot plant; its main objectives to establish basic operating parameters and generate material for initial environmental testing. Prior to the five run-in pilot plant runs, batch flotation tests were conducted on the available composites.

The fourth phase of metallurgical testing was a pilot plant campaign carried out at Lakefield Research Limited in August, 1997 under the supervision of Melis (Melis Project No. 345, November 27, 1998 Report). This phase included batch and locked cycle tests as well as pilot plant runs carried out on composites prepared from assay rejects and half core representing different zones of the Prosperity deposit. It also included grinding test work, detailed analysis of concentrates, generation of environmental data, tailings settling tests, and concentrate settling and filtration tests.

A test work program was initiated in 2008 to investigate the metallurgical performance of Prosperity mineralization at coarser regrinds and alternate regrind flow sheets. This work was conducted at Process Research Associates in Richmond, British Columbia on core from the 2007 drilling campaign that represented years 1 through 4 mill feed.

16.2 Composites

Phase II Composite Preparation

A sketch of the diamond drill hole plan used for the gathering of metallurgical samples for the Phase II test program is shown in Figure 16-1.

Individual composites were prepared by Min-En Laboratories of Vancouver, British Columbia. A total of 24 individual composites were made up from assay rejects using a 1 kg weight for each m of intersection.

Figure 16-1

Phase II Test Program — Drill Hole Plan

Phase III Composite Preparation

Approximately 21 tonnes of samples were received for run-in pilot plant testing, including both assay rejects and drill core from the 1994 drilling program representing the upper level (>200 m) of the main (east) area of the mineralization. Assay rejects were blended as separate intrusive and volcanic sub-composites for batch testing and one overall composite for run-in pilot plant runs RI-PP1, RI-PP2 and RI-PP3. Drill core samples were blended as separate intrusive and volcanic sub-composites for batch testing and one overall composite for run-in pilot plant runs RI-PP4 and RI-PP5.

Two kilogram test charges were prepared for five alteration composites (intrusive sericite, intrusive potassium silicate, volcanic sericite, volcanic potassium silicate and volcanic propylitic) for use in batch flotation tests to generate environmental data.

A sketch of the diamond drill hole plan of the main zone used in the 1994 metallurgical sampling for the Phase III metallurgical test work is shown in Figure 16-2.

Figure 16-2 Phase III Test Program — Drill Hole Plan

Phase IV Composite Preparation

Test composites were prepared for pilot plant testing from 58 tonnes of assay rejects and half core samples. The assay rejects samples were blended into three composites (upper, middle, and lower) for batch testing and one overall composite for locked cycle and pilot plant testing. Half core samples were blended into three composites (upper, middle and lower), which were run separately in batch tests, locked cycle tests, and pilot plant runs.

A sketch of the drill hole plan for the 1996/1997 drilling program (Phase IV), showing the locations of the drill holes in the Prosperity deposit, is shown in Figure 16-3.

Figure 16-3

Phase IV Test Program — Drill Hole Plan

16.3 Mineralogy

The major copper mineral in the Prosperity deposit is chalcopyrite (CuFeS2). It occurs in multistage veins and associated fracture-controlled gangue segregation. In veins the metallic assemblage mostly includes chalcopyrite-pyrite (FeS2) with some local occurrence of bornite (Cu5FeS4) -tetrahedrite [(Cu,Fe,Zn,Ag)12Sb4S13)]/tennantite [(Cu,Fe,Zn,Ag)12As4S13)] and minor occurrences of chalcocite (Cu2S) or trace digenite (Cu9S5). Local occurrence of molybdenite (MoS2) and gold also occur. In vein-related segregation chalcopyrite and pyrite are the main metallic minerals with minor occurrences of bornite.

Altered - impregnated {quartz(SiO2) — sericite[KAl2(AlSi3O10)/(OH)2] — albite (NaAlSi3O8)0.9(CaA12Si2O8)0.1} groundmass contains very minor finely disseminated chalcopyrite as loose clusters. Disseminated pyrite occurs abundantly in quartz-sericite altered rocks in a biotite [K(Mg,Fe)3AlSiO10(OH)2] chlorite [(Mg,Al,Fe)12(Si,Al)8O20(OH)16] hornblende and chlorite-carbonate [calcite(CaCO3)/dolomite(CaMg(CO3)2)] altered mafics. Significant amounts of chalcopyrite occur interstitially to and as small blebs in magnetite (FeFe23+O4) — hematite (Fe2O3).

Chalcopyrite, with local occurrences of bornite, tetrahedrite/tennnantite and lesser chalcocite, is commonly associated with pyrite as interstitial microveinlets and as scattered small inclusions in pyrite.

Native gold shows strong zonal distribution and commonly occurs in loose clusters of grains with strong vein control in quartz and carbonate gangue. Size ranges noted were <2.5 µm (microns) to 95 µm. It occurs most abundantly in association with copper minerals, particularly where tetrahedrite/tennantite and chalcocite are present. It was noted as isolated grains in gangue, as isolated blebs and microveinlets in pyrite, as blebs in chalcopyrite, tetrahedrite/tennantite and chalcocite, and rarely with bornite.

Molybdenite has zonal distribution and occurs as isolated grains and loose clusters of grains. It is vein/segregation controlled and is associated with chalcopyrite in quartz/sericite — chlorite gangue.

Non-opaque gangue materials include quartz, sericite and feldspar (K, Na, AlSi3O8) with subordinate amounts of carbonate (CO3). Opaque gangue minerals include mainly pyrite with subordinate amounts of iron or titanium oxides and iron hydrides (magnetite, hematite, rutile (TiO2), goethite [FeO(OH)].

Gypsum (CaSO42H2O) and lesser anhydride (CaSO4) form late veins and open space fillings in and crossing mineralized veins, and form-filling voids in gangue segregation. There is a spatial relationship between gypsum (anhydride) and mineralization with accidental inclusion of sulphides in gypsum.

16.4 Grinding

Work Indices

The abrasion indices, bond impact crushing work indices, correlated autogenous work indices, rod mill bond work indices, and ball mill bond work indices measured for the Prosperity composites in the three test programs are listed in Table 16—1.

Table 16-1

Metric Work Indices

		Bond Impact
		Crushing	Correlated	Rod Mill
	Abrasion	Work	Autogenous	Bond	Ball Mill
Composite	Index (1)	Index	Work Index	Work Index	Work Index
Sept/91	—	—	21.7	17.4	15.9
U1	—	—	18.2	20.6	16.8
U2	—	—	16.5	19.8	15.3
M1	—	—	17.0	17.7	15.9
Int.	—	—	—	15.0	15.5
Vol.	—	—	—	16.4	16.1
VU(AGL)	0.0952	5.5	18.3	16.1	17.7
VU(BGL)	0.1832	3.3	—	16.1	17.7
VM	0.1444	7.5	20.0	17.3	17.9
VL	0.3421	5.0	20.2	18.6	20.4
IU(AGL)	0.1339	6.7	—	16.2	16.4
IU(BGL)	0.2181	—	—	16.2	16.4
IM	0.2638	5.4	—	15.5	17.5
IL	0.3660	5.8	—	14.7	17.6
Averages:
All	0.2183	5.5	18.8	17.1	16.9
Volcanic	0.1912	5.3	19.5		18.0
Intrusive	0.2455	5.7	—		16.8
Upper	0.1576	5.1	17.7		16.6
Middle	0.2041	6.5	18.5		17.1
Lower	0.3541	5.4	20.2		19.0

On average the Prosperity mineralization has a low abrasion index and a relatively low impact crushing index. The measured grinding indices showed the mineralization to be of medium hardness, but with a relatively high autogenous work index.

Grinding Tests

Three batch tests were completed on the Overall Half Core Composite to analyze the effect of differing primary grind product discharge sizes on copper and gold recoveries.

Based on these three tests, a cost benefit analysis using the value of incremental copper and gold recoveries against net grinding power costs concluded that the optimum 80% cumulative passing size (K80) of the primary grind was approximately 160 µm.

Four batch tests were completed to analyze the effects of regrind size distribution on copper and gold recovery to the third cleaner concentrate.

The optimum regrind K80 appears to be between 14 µm and 17 µm in the regrind product. Gold recovery was found to be more sensitive to the K80 of the regrind product than copper recovery.

16.5 Gravity Separation

A total of five gravity recovery tests were conducted during the Phase III metallurgical test work to determine the potential for gold recovery by gravity separation. Gravity gold recovery was performed on drill core composites ground to match the flotation test grind. Processes tested were a Wilfley Table followed by a Mozley Concentrator, a Falcon Separator followed by a Mozley Concentrator, a Knelson Concentrator followed by a Mozley Concentrator, and a Falcon Separator.

Gravity gold recovery on the intrusive composite was 6.1% to 9.5%, into a relatively low grade upgraded concentrate (153 to 269 gpt Au). A single test on copper scavenger tails, although not definitive, showed there may be potential to recover gold from the copper scavenger tails by gravity.

16.6 Batch Tests

Phase II

Batch flotation tests were carried out on composites and sub-composites of the Prosperity mineralization in each of the second, third and fourth phases of metallurgical test work.

The second phase batch tests were in two categories; development tests and variability tests. Development tests investigated recovery processes to determine those best suited to the Prosperity mineralization.

Once an acceptable flotation scheme had been identified, those conditions were used in a series of variability batch tests to obtain a measure of metallurgical variability across the deposit for the upper, middle and lower zones.

Acceptable copper and gold recoveries were achieved in these tests with cleaner concentrate grades generally being in the range of 20% Cu to 30% Cu and approximately 45 gpt Au.

Phase III

Batch flotation tests were conducted in the third phase of metallurgical test work to determine if intrusive and volcanic drill core composites could be blended for run-in pilot plant testing. The batch tests yielded good bulk recoveries (up to 95.7% copper recovery and 79.9% gold recovery for the intrusive composite; 93.4% copper recovery and 82.4% gold recovery for the volcanic composite). Good cleaner concentrate grades were also achieved (up to 25.1% Cu and 41.9 gpt Au for the intrusive composite; 28.6% and 51.6 gpt Au for the volcanic composite). Other than head grade differences there were no metallurgical differences between the intrusive and volcanic rock types.

Phase IV

In the fourth phase of metallurgical test work, batch flotation tests were completed on volcanic and intrusive sub-composites, and on upper, middle, and lower composites prepared from assay rejects and half core composites.

Good results were achieved with respect to both copper and gold metallurgy using a relatively coarse primary grind and a conventional bulk rougher/copper-pyrite separation float. Batch tests completed on the assay rejects composites yielded poorer copper recoveries than batch tests completed on the half core composites, and when blended with half core composite the assay rejects composite lowered the copper recovery.

From the results of these tests it was decided to mix the assay rejects samples into a single composite to run in the pilot plant, and mix separate upper, middle and lower half core composites for definitive pilot plant runs.

16.7 Locked Cycle Tests

Locked cycle tests were carried out on composites and sub-composites of the Prosperity mineralization in Phase II and Phase IV of the metallurgical test work.

In the second phase of metallurgical test work locked cycle tests were conducted on 11 composites to investigate the metallurgical performance of the Prosperity mineralization under conditions approaching steady state.

These locked cycle test results were used to provide estimates of copper and gold recoveries for mine block model development. The average results were 85.6% copper recovery and 72.1% gold recovery into a concentrate grading 25.2% Cu and 46.2 gpt Au from an average head grade of 0.22% Cu and 0.47 gpt Au.

In the fourth phase of metallurgical test work, eight locked cycle tests were conducted prior to and following the pilot plant runs to confirm conditions for the pilot plant runs and provide data to increase the level of confidence regarding estimates of copper and gold recoveries. The locked cycle tests completed in Phase IV were conducted using test conditions reviewed and revised during the Phase IV batch tests.

The copper results for the eight-cycle locked cycle tests (average copper recovery of 88.7%) were slightly inferior to those from the six-cycle locked cycle tests (average copper recovery of 90.8%), but comparisons of bulk rougher recoveries show that the difference was negligible. The gold results from the eight-cycle locked cycle tests (average gold recovery of 77.1%) were superior to those from the six-cycle locked cycle tests (average gold recovery of 73.8%), both in comparison of the overall recoveries and bulk rougher recoveries. Concentrate grades achieved for the eight-cycle locked cycle tests, which averaged 26.3% Cu and 44.9 gpt Au from an average head grade of 0.22% Cu and 0.44 gpt Au, were also superior to those achieved in the six-cycle locked cycle tests, which had an average concentrate grade of 22.4% Cu and 36.6 gpt Au from an average head grade of 0.22% Cu and 0.45 gpt Au.

16.8 Run-In Pilot Plant and Main Pilot Plant Runs

The run-in pilot plant was conducted in the third phase of the metallurgical test work. Five runs were completed, three runs on a composite of volcanic assay rejects and the final two runs on combined volcanic and intrusive drill core composites. All composites represented the upper zone (<200 m) of the main (east) area of the deposit.

Because of limited sample availability, only short pilot plant runs were possible and stable circuit conditions, especially for gold, were not reached. Consequently, generally inferior results were achieved. In spite of this, run RI-PP5 bulk rougher flotation conditions yielded good bulk copper recovery (92.1%) and acceptable (79.5%) gold recovery, even at a relatively coarse primary grind K80 of 157 µm.

Despite the slightly oxidized nature of the drill core composites in run RI-PP4, it was possible to achieve a respectable copper concentrate grade of 22.4% Cu and 21.5 gpt Au. The main pilot plant was conducted in Phase IV of the metallurgical test work. Eight runs were completed, five runs on an assay rejects composite and three runs on half core composites.

The average copper recovery increased from the 82.0% recovery achieved in the run-in pilot plant to an average of 86.2% for the two pilot plant runs on the assay rejects composite in which stable steady-state operating conditions were achieved and 87.6% for the runs on the half core composites. The gold recovery, which averaged 55.6% for the run-in pilot plant increased to 79.0% for the runs on the assay rejects composite but only averaged 66.5% for the runs on half core composites.

The results of the two stable pilot plant runs on the assay rejects composite (average recovery of 86.2% for copper and 79.0% for gold) were in close agreement with the six-cycle locked cycle test result (85.6% copper recovery and 78.3% gold recovery).

The pilot plant copper recovery for the half core composites (average of 87.6%) were slightly inferior to the six-cycle locked cycle test copper recoveries (average of 90.8%). In terms of gold recovery, the upper, middle and lower pilot plant recoveries (60.4%, 73.6% and 65.5% respectively) were lower than the respective six-cycle locked cycle test recoveries (68.1%, 75.2% and 78.2%).

It was believed that an overly coarse regrind cyclone overflow in the pilot plant runs on the half core composites was the cause of the low gold recoveries. Batch and eight-cycle locked cycle tests completed after the pilot plant confirmed this, when excellent copper and gold recoveries (average of 88.7% for copper and 77.1% for gold) and concentrate grades (average of 26.3% Cu and 44.9 g/t Au) were achieved on the upper, middle and lower half core composites.

16.9 Target Concentrate Grades and Recoveries

The locked cycle tests from Phase II and Phase IV of the metallurgical test work were used to derive target concentrate copper and gold recoveries and copper concentrate grades.

Target copper and gold concentrate grades and recoveries for two different mill feed grades (Cases A and B) are estimated in Table 16-2.

Table 16-2

Target Gold & Copper Recoveries and Concentrate Grades

				Concentrate
	Case	Head Grade		Grade		% Recovery
Zone	No.	% Cu	gpt Au	% Cu	gpt Au	Copper	Gold
Upper				25.1	36.1	90.1	70.4
Middle/Lower	A	0.236	0.434	25.1	39.5	90.1	77.1
Upper	B	0.246	0.482	25.5	40.0	90.2	72.2
Middle Lower				25.5	43.8	90.2	79.0

Calculations to estimate copper recovery based on copper head grade were derived as shown in Table 16-3.

Table 16-3

Target Copper Recovery & Target Concentrate Copper Grade Calculations:

Lower, Middle and Upper Zones

Head Grade Range		% Cu in
(% Cu)	% Copper Recovery	Concentrate
0.10 to 0.20	(% Cu x 93.0) + 71.3	(% Cu x 48.0) + 14.2
0.20 to 0.25	(% Cu x 6.0) + 88.7	(% Cu x 36.0) + 16.6
0.25 to 0.40	(% Cu x 6.0) + 88.7	(% Cu x 6.7) + 23.9
>0.40	91.0	26.6

Target gold recovery calculations, to estimate gold recovery based on gold head grade, were derived as show in Table 16-4 and 16-5.

Table 16-4

Target Gold Recovery Calculations

Upper Zone

Head Grade Range
(gpt Au)		% Gold Recovery
0.10 to 0.40		(gpt Au x 97.0) + 30.3
0.40 to 0.65		(gpt Au x 38.0) + 53.9
0.65 to 1.00		(gpt Au x 5.7) + 74.9
>1.00		81.0

Table 16-5

Target Gold Recovery Calculations

Middle & Lower Zones

Head Grade Range (gpt Au)		% Gold Recovery
0.10 to 0.29		(gpt Au x 110.0) + 39.4
0.29 to 0.49		(gpt Au x 40.0) + 59.7
0.49 to 1.00		(gpt Au x 7.5) + 75.6
>1.00		83.0

The target concentrate gold grade can be calculated from the copper head grade, copper recovery, concentrate copper grade, gold head grade, and gold recovery as follows:

Target Concentrate Gold Grade (gpt Au) =

(Concentrate Copper Grade, %) x (Gold head grade, gpt Au) x (Gold Recovery, %)

(Copper head grade, %) x (Copper Recovery, %)

Figure 16-4 displays the target copper recovery versus copper head grade.

Figure 16-4

Target Copper Recovery Vs Copper Head Grade

Figure 16-5 displays the target concentrate copper grade versus copper head grade:

Figure 16-5

Target Concentrate Copper Grade Vs Copper Head Grade

Figure 16-6 displays the target gold recovery versus gold head grade:

Figure 16-6

Target Gold Recovery Vs Gold Head Grade

16.10 Concentrate Analysis

An assessment of concentrate analyses from the locked cycle tests and pilot plant runs carried out in Phase IV of metallurgical testing yielded the values shown in Table 16-6 as typical concentrate analysis obtained in the Prosperity pilot plant program:

Table 16-6

Typical Concentrate Analysis

					ppm
Zone	% Cu	gpt Au	gpt Ag	% As	Hg	% Pb	% Sb	% Zn
Upper	23.1	42.8	104	0.25	144	0.36	0.38	3.57
Middle	24.4	40.1	91	0.27	105	0.19	0.37	1.02
Lower	25.3	40.1	81	0.14	68	0.06	0.29	0.55

Settling tests were completed on the bulk final concentrate from the Phase IV pilot plant run, the run carried out on the upper core composite. Results are summarized in Table 16-7.

Table 16-7

Settling Tests on Concentrate from Upper Composite

			Density [%
	Percol 351 Addition		Solids(w/w)]		Supernatant	Concentration Zone
Test		Liquid			Clarity	Settling Rate	Unit Area
No.	Solids (gpt)	(mg/L)	Initial	Final	(ppm)	(m3/m2/d)	(m2/t/d)
S9	20.3	4.9	19.6	69.6	62	319.1	0.084
S10	11.3 + 5.6	4.1	19.7	71.0	2	326.4	0.084
S11	0	0	19.8	75.3	80	29.4	0.299

Pressure filtration testing by Larox Inc. achieved concentrate moistures of 8.0% to 9.1% solids (w/w). Filtration rates as high as 544 kg/m2/h were obtained.

CERAMEC filtration tests conducted by Outokumpu Minetec USA Inc. achieved a moisture level of 14.5 to 14.8% solids (w/w) with filtration rates of 185 kg/m2/h to 327 kg m2/h.

16.11 Tailings Settling Tests

Four sets of flocculent scoping tests were completed on the combined tailings produced in Locked Cycle Test No. M2 in Phase II of metallurgical testing. These tests indicated that the preferred flocculent was Percol 919, and that an addition rate of 20 to 25 gpt was satisfactory.

Following the scoping tests, fifteen variability settling tests were conducted on tailings generated in the six locked cycle tests completed on the west zone and main zone composites.

In Phase III, a series of four settling tests were performed on run-in pilot plant run RI-PP5 tailings.

In Phase IV of metallurgical testing, separate settling tests were performed on each final tailings stream from the three half core pilot plant runs. Settled tailings densities of 62 to 66% solids (w/w) were achieved with thickener unit areas varying from 0.247 to 0.351 m2/t/d

16.12 Environmental Data

Tails solids and liquid from three Phase II locked cycle tests on composites were collected and submitted to Saskatchewan Research Council (Saskatoon, Saskatchewan) for detailed low-level analysis. Tails from Phase II locked cycle tests were submitted for detailed environmental analysis. Data collected included detailed analysis of tails solids, tails liquid (total and dissolved), standard Special Waste Extraction Procedure (SWEP), simulated rainfall test, acid-base accounting, and bio-assay tests.

In Phase III, environmental data was collected for tails samples from run-in pilot plant run RI-PP5 and from five alteration composites (intrusive sericite, intrusive potassium silicate, volcanic sericite, volcanic potassium silicate, and volcanic propylitic). Phase IV environmental data was collected from tails samples from pilot plant runs PP6, PP7 and PP8. Data collected included detailed analysis of head samples, tailings solids and tailings liquid, acid-base accounting measurements on heads and tailings samples, rainbow trout and daphnia magna toxicity tests on tailings decant liquid, and separate tailings aging tests with analysis of solids and liquids for the run-in pilot plant run RI-PP5, and the pilot plant runs on the upper, middle, and lower half core composites.

Presentation of this data has been provided in the Melis Phase II, Phase III and Phase IV reports. Analyses of the environmental data have been conducted by environmental experts.

16.13 Regrind Evaluation Test Work

Test work was conducted at a variety of regrind sizes and confirms earlier work that indicated improvements in copper and gold concentrate grades and recoveries with decreasing particle size. A regrind 80 % passing size of 17 um was selected due to performance consistent with the project economics. Limited test work at a regrind 80 % passing size of 11 um indicated that there may be opportunity for recovery and grade improvements at these ultrafine regrinds.

Test work exploring the impact of a split regrind flow sheet, where the initial concentrate from rougher flotation was ground to a series of targets coarser than the 17 um target and the remaining concentrate was ground to the 17 um target was conducted. This strategy yielded results similar to the single regrind flow sheet only when additional cleaner flotation time was added.

Given these results a 17 um regrind design target and flow sheet was carried forward.

17. Mineral Resource and Mineral Reserve Estimates

17.1 Resource Modeling

The property exploration and sampling work culminated in the comprehensive geological interpretation and delineation of a major, porphyry-style gold-copper resource, based on a total of 150,090 m of core drilling in 402 holes and 6,309 m of drilling in 68 percussion holes, with an average drill hole spacing of 70 m.

In January 1998 Giroux Consultants Ltd. (GCL) commenced geostatistical analysis of data from the Prosperity Project Database. The Prosperity Project database consists of all relevant drill hole collar and downhole survey data, assay and ICP data, and geological and geotechnical data.

Capping and Compositing of Assay Results

An analysis of high-grade outlier values was conducted in order to determine a ‘capping’ methodology prior to compositing of assays. Respecting geological boundaries, GCL created 8,976 downhole composites 15 m in length. A total of 8 distinct geologic domains were coded in preparation for variography and ordinary kriging of block grades. An additional domain code was assigned solely to the post-mineral dikes large enough to form discrete blocks (irrespective of alteration or geographic domain). These blocks were not kriged but a global average grade was assigned, as discussed in the section on post mineral dikes below. These composites were used to estimate the gold and copper grades of 289,638 model blocks. The dimensions of the blocks are 20 m by 20 m horizontally and 15 m thick.

Variography

GCL calculated a series of semi-variograms for geologic domains for which there was sufficient data. For domains with insufficient composite data, simplified spherical models were fitted. For modeling purposes the various alteration, lithologic, and geographic domain information was combined into a number of unified ‘geologic’ domains to which the assay data could be linked. Within this block model, Cu and Au grades for each block were estimated independently by ordinary kriging, utilizing search parameters obtained from the variography analyses.

Kriging

Using the domains established above, a grade model was estimated by ordinary kriging within a three-dimensional grid comprising 20 x 20 x 15 m blocks coded by geologic domain. The block model extents in Mine Grid are shown in Table 17.1.

Table 17-1

Block Model Extents

	Easting	Northing	Elevation
Lower SW corner	8990E	8610N	540
Upper NE corner	11510E	11010N	1605
Number of blocks	126	120	71

Within the block model, Cu and Au grades for each block were estimated independently by ordinary kriging, using the search parameters obtained from the variography. Initial search ellipsoids were defined as 1/3 of the variogram range. Provided that a minimum of four composites were found within the initial search ellipsoid for the appropriate element, a grade estimate was calculated. However, if the minimum four composites were not found in the initial search, the axes of the ellipsoids were expanded by a factor of 1.5. If the requisite composites were still not found the expansion was then doubled to the full range of the variogram. If after two expansions, the minimum four composites were still not found, then the block value was not estimated. Alternatively, if more than 15 composites were found during the search procedure, the 15 closest to the block centroid were used for estimation.

From previous work it was known that a number of sporadic, narrow, high-grade gold veins existed, particularly in the propylitic rocks in the periphery of the deposit. In order to prevent these values from unduly influencing large volumes of surrounding rock it was necessary to restrict their influence. The procedure used was as follows: individual assays in propylitic rocks with gold grades in excess of 1 g/t were identified, and the corresponding composites flagged (flag = 2). If a search ellipsoid located a gold vein composite, within the block to be estimated, the composite was used in the estimation of that block. If a search ellipsoid located a flagged gold vein composite which was outside the block to be estimated, it was not used to estimate that block. A similar strategy was adopted for dealing with the narrow, low grade post mineral dikes, which essentially have the opposite effect on grade estimation of nearby blocks as the high-grade veins.

17.2 Resource Classification

Each block within the model was classified as measured, indicated or inferred in terms of resource classification, corresponding with proven or probable respectively, in terms of reserve classification. A kriging estimation error was calculated for each estimated block for both copper and gold which takes into account the nugget effect, sill value, number of composites used in the estimate and the spatial relationships of composites relative to any anisotropy. The kriging estimation error was then used to compute a grade-linked relative kriging estimation error as follows:

Relative Estimation Error = (Kriging Estimation Error / grade) * 100%

For both Cu and Au, separate histograms of these relative estimation errors were plotted and the data divided into three populations based on the magnitude of the error. With reference to these three divisions, resource blocks were classified for each element as measured, indicated, inferred according to the set of rules summarized below:

Measured

· Blocks with relative estimation errors for Cu of less than 24%

· Blocks with relative estimation errors for Au of less than 40%.

Indicated

· Blocks with relative estimation errors for Cu greater than or equal to 24% and less than 49%

· Blocks with relative estimation errors for Au greater than or equal to 40% and less than 70%.

Inferred

· Blocks with relative estimation errors for Cu greater than or equal to 49%.

· Blocks with relative estimation errors for Au greater than or equal to 70%

Since the economics of the deposit depend on both gold and copper, the final classification of the (in-situ) resource block considered the combined effects of the individual block classifications. This, in turn necessitated an additional set of rules, as set out below:

· For the overall block to be classed as measured, relative estimation errors for both Cu and Au had to fall in the measured category.

· If relative estimation errors for either Cu or Au fell into the indicated category, while the other metal was classed as measured, then the overall block was classed as indicated/probable.

· If relative estimation errors for both Cu and Au fell into the indicated category then the overall block was classed as indicated.

· If relative estimation errors for either Cu or Au fell into the inferred category, then the overall block was classed as inferred, irrespective of the classification of the other value.

· If during kriging, expansion of the initial search ellipsoid of 1/3 the variogram range was required to estimate either copper or gold, then the overall block was classed as inferred.

· Any blocks assigned a grade directly (e.g. blocks of post mineral dike) were classed as inferred.

In December 1998 the classification scheme originally applied by Giroux was re-examined, for the reasons outlined below. In particular the use of the expanded search criteria was reevaluated. In discussions with Giroux, it was mutually agreed that within the core of the deposit as defined by the proposed ultimate pit at the time, the relative estimation error alone was sufficient for classifying the mineral resource. However, outside this boundary, the expanded search criteria in addition to the relative estimation error were applied in classifying resources.

The following is a brief clarification on the classification methodology. During estimation of the resource model, the expanded search criterion of 1/3 the variogram range was used to protect against over-classifying blocks on the margins of the deposit, which could only be estimated using an expanded search owing to a paucity of drill information and relatively long variogram ranges. The outcome of this however, was that certain blocks in the core of the deposit, that met all the other requirements for being classed as measured and/or indicated, were classed as inferred, simply because they required an expanded search during kriging. A number of these blocks even had data points within their boundaries. To accommodate this problem during classification of resources within the main deposit area, the practice of classifying every block estimated during an expanded search as inferred was dropped, and blocks within the optimized pit were re-classified, solely on the basis of relative kriging estimation errors.

Table 17-2

Prosperity Mineral Resources

at 0.14% Copper Cut-off

	Tonnes	Gold	Copper
Category	(millions)	(gpt)	(%)
Measured	547.1	0.46	0.27
Indicated	463.4	0.34	0.21
Total	1,010.5	0.41	0.24

17.3 Reserve Estimation

Pit Optimization

Historically, the pit optimization was done by developing nested pit shells by varying discount rates. Schedules and cashflows were developed for each of those shells at varying concentrator throughputs and capital costs. Analysis resulted in the selection of a 70,000 tpd processing rate and a design basis shell that optimized IRR and NPV.

In 2009 the throughput rate of 70,000 tpd was re-evaluated based on current capital estimates and long term metal prices. The results were consistent with previous work.

In 2009, higher metal prices triggered a re-evaluation of the Pit Optimization. The work was performed by AKF Mining Services Inc. using GEMCOM-WHITTLE Optimization software (based on the Lerchs-Grossman (LG) Algorithm).

The software generates a series of undiscounted break-even nested pit shells from a range of copper and gold prices. The metal price range was derived by applying a range of revenue factors to base copper and gold prices of US$1.25/lbs and US$500/oz, respectively. Based on a 70,000 tpd throughput rate, the discounted cashflow for each shell determines a maximum economic pit limit for each price. The results of the cashflow analysis were used to guide the optimum design limits to satisfy this study.

Parameters used in Optimization

Table 17.3 summarizes the parameters used in the optimization. All costs are in Canadian Dollars unless otherwise noted.

Table 17-3

Pit Optimization Parameters

Parameters		Optimization Input
Cu Price		US$1.25/lb (US$0.98/lb not including offsite costs)
Au Price		US$500/oz (US$380/oz not including offsite costs)
Exchange Rate		US$0.74/CDN$1.00
Revenue Factor (RF)		0.5 to 2.0
Mining Cost		$1.20/tonne of Material
Bench Increment		$0.03/tonne/bench — 1485 Elev. Entrance
Milling + G&A Cost		$4.20/tonne of Ore
Cu & Au Recovery		90% & 70%
Cashflow Discount Rate		10%
Overall Slopes		W & NE Sector = 40°, Else = 42°

A capital of $850 million was used in the optimization cash flow analysis and the revenue parameters were only applied to measured and indicated ore blocks.

Wall Slopes

Pit wall slope recommendations were provided in a report by Knight Piesold, “2007 Feasibility Pit Slope Design, dated September 21, 2007.

For the pit optimization, the maximum pit depth was estimated at 850 meters and the overall slope was adjusted for 3 ramp crossing, hence the estimated slopes for West and Northeast sectors is 40° and for the remaining sectors is 42°.

Pit Selection

The line graphs in Figure 17-1 represent the best and worst case value scenarios for each shell. The best case graph plots discounted values based on mining performed shell by shell. The worst case graph plots discounted values based on mining performed bench by bench. This gives a representation of where the optimized pit shell lies for each best and worst case curve.

Figure 17-1 GEMCOM-WHITTLE Pit by Pit Graph

The selection of a pit shell also considers discounted value, mill throughput, life of project and risk. Analyzing figure 17-1, Pit 10 (RF = 0.95) has the maximum discounted value for the best case assumptions, and Pit 4 (RF = 0.65) has the maximum discounted value for the worst case assumptions. Between Pit 10 and 31, the best case curve shows the incremental value is minimal and for the worst case the incremental value is decreasing. Since the selection of optimum and intermediate shells will be based on the best case scenario curve, these analyses conclude that Pit 10 satisfies the criteria for this study.

Pit Design

Using Pit 10 from the optimization work, the shell was imported into Minesight software, to perform the detailed design work, which is referenced in detail in section 18-2.

NSR Pit Rim Cut-off

An NSR calculation was then applied to the block model and a CDN$5.50/tonne NSR pit rim cutoff used to report the tonnes and grades for the current mineral reserves. The basis for the $5.50/tonne NSR pit rim cutoff is shown in Table 17-4.

Table 17-4 NSR Pit Rim Cut-off Calculation

Milling		$3.85/tonne ore
G&A		$0.52/tonne ore
Rehandle Allowance		$0.65/tonne ore
Sustaining Capital & Margin		$0.83/tonne ore
NSR Pit Rim Cut-off		$5.50/tonne ore

Net Smelter Return Model

The Net Smelter Return (NSR) was updated for the block model to calculate insitu ore value based upon metal content, metallurgical recovery, metal price, off property costs and currency exchange rate. The model then provides an estimate of the NSR value for each block expressed in CDN$/t for all interpolated blocks in the geological resource model. The NSR value incorporates the following economic evaluation criteria:

· Variable metallurgical recovery and concentrate grade based upon head grade for copper and gold in the Upper, Middle and Lower geological domains

· A fixed silver concentrate grade based upon test work results.

· Concentrate transportation, treatment, refining, and penalties

· Fixed metal price and exchange rate parameters

An example of the NSR calculation is shown in Table 17-5. The metallurgical recoveries were assigned first and a gross metal value was calculated. Adjustments were then made for transportation, treatment and refining charges. The NSR value was then calculated for each block of the resource model and stored in the model.

Table 17-5 2009 Net Smelter Return Calculations

Parameters and Test Block Calculation

Test Block NSR Calculation
Copper Head Grade	%															0.218
Gold Head Grade	g/t															0.427
Location																Lower Zone
Metallurgical Recovery
Copper Recovery
		Cu%						Upper, Middle & Lower Copper Recovery								%
Head Grade Range	>=	0.100			<			0.200		Cu x		93.0	+	71.3	=	91.57
Head Grade Range	>=	0.200			<			0.400		Cu x		6.0	+	88.7	=	90.01
Head Grade Range	>=	0.400													=	91.00
Gold Recovery
		Au g/t						Upper Zone Gold Recovery								%
Head Grade Range	>=	0.100			<			0.400		Au x		97.0	+	30.3	=
Head Grade Range	>=	0.400			<			0.650		Au x		38.0	+	53.9	=
Head Grade Range	>=	0.650			<			1.000		Au x		5.7	+	74.9	=
Head Grade Range	>=	1.000													=	81.00
		Au g/t						Lower & Middle Zone Gold Recovery								%
Head Grade Range	>=	0.100			<			0.290		Au x		110.0	+	39.4	=	86.37
Head Grade Range	>=	0.290			<			0.490		Au x		40.0	+	59.7	=	76.78
Head Grade Range	>=	0.490			<			1.000		Au x		7.5	+	75.6	=	78.80
Head Grade Range	>=	1.000													=	83.00
Metal Pricing
Copper Price	US$/lb	$	1.65													$	1.65
Gold Price	US$/ounce	$	650.00													$	650.00
Silver Price	US$/ounce	$	10.00													$	10.00
USD:CDN Exchange		0.82														$	0.82
Concentrate
Copper Concentrate Grade
					% Cu			Copper Concentrate Grade								Cu%
Head Grade Range	>=	0.100			<			0.200		Cu x		48.00	+	14.2	=	24.66
Head Grade Range	>=	0.200			<			0.250		Cu x		36.00	+	16.6	=	24.45
Head Grade Range	>=	0.250			<			0.400		Cu x		6.70	+	23.9	=	25.36
Head Grade Range	>=	0.400													=	26.60
Gold Concentrate Grade	g/dmt	Calculated from head grade, recovery and concentrate production
Silver Concentrate Grade	g/dmt	89.00														89.00
Moisture Content	%	8.0		%												8.0		%
Contained Copper	lb/dmt															538.83
Contained Gold	g/dmt															40.85
Contained Silver	g/dmt															89.00
Payable Copper	lb/dmt															516.79
Payable Gold	g/dmt															39.83
Payable Silver	g/dmt															84.55
Concentrate - Recovery Based	dmt/t ore															0.00803
Gross Value of Concentrate after Deductions
Gross Value Concentrate	C$/dmt															$	2,088.08
Concentrate Handling
Truck Haul Mine to Rail	C$/wmt	$	28.00
Rail Freight	C$/wmt	$	23.00
Stevedoring	C$/wmt	$	24.50
Ocean Freight	US$/wmt	$	60.00
Total Concentrate Handling	C$/wmt	$	148.67
	C$/dmt	$	161.60													$	161.60
Treatment and Refining
Deduction for Copper	unit	1.00			95.9		%									1.00
Treatment Charges	US$/dmt	$	80.00													$	80.00
Gold Payment	%	97.5%														97.5		%
Silver Payment	%	95.0%														95.0		%
Copper Refining Cost	US$/payable lb	$	0.08													$	0.08
Gold Refining Cost	US$/payable oz	$	6.00													$	6.00
Silver Refining Cost	US$/payable oz	$	0.45													$	0.45
Total Treatment and Refining	C$/dmt															$	158.84
		As			Sb			Hg		Total
Penalties Upper Zone	US$/dmt	$	4.50		$	8.40		$	24.80	$	37.70			C$/dmt		$	45.98
Penalties Middle Zone	US$/dmt	$	5.10		$	8.10		$	17.00	$	30.20			C$/dmt		$	36.83
Penalties Lower Zone	US$/dmt	$	1.20		$	5.70		$	9.60	$	16.50			C$/dmt		$	20.12
Net Smelter Return
Net Smelter Return	C$/dmt															$	1,747.52
	C$/payable lb Cu															$	3.38
NSR	NSR $/t															$	14.03

Metallurgical Recovery

The metallurgical recovery relationships used in the NSR calculation were based upon results of the 1997 Pilot Plant program conducted at Lakefield Research under the direction of Melis Engineering. As part of the resource modeling process the deposit was subdivided laterally and vertically into three geological domains. For the purposes of NSR calculation the Upper and Middle/Lower Domains were assumed to have quantifiable differences in gold recovery.

Metal Prices

The copper concentrate will contain three payable metals — copper, gold and silver. The NSR calculation was based upon the following metal prices:

	Copper	US$1.65/lb
	Gold	US$650/ounce
	Silver	US$10.00/ounce

The exchange rate used for converting US$ to CDN$ was US$0.82/CDN$ 1.00.

Concentrate Grade

The copper concentrate grade for copper was calculated for each block based upon the copper grade. Four copper grade ranges were established with different copper concentrate grade formulae. Blocks below 0.1% Cu were not considered and blocks above 0.4% Cu were assigned a maximum concentrate grade of 26.6% Cu.

The gold concentrate grade was based upon the gold head grade, estimated recovery and the copper concentrate production from the theoretical block grade.

The silver concentrate grade was fixed at 89 grams per dry metric tonne (g/dmt).

Concentrate Penalties

Concentrate penalties for arsenic, antimony and mercury were calculated based upon the location of the block in the upper, middle or lower geological domains.

Transportation Cost

The concentrate transportation costs include trucking, rail haulage, stevedoring, and ocean freight. These were identified on a wet metric tonne basis and summarized for 8% moisture content in Canadian dollars per dry metric tonne of concentrate.

18. Additional Requirements for Technical Reports on Development Properties and Production Properties

18.1 Site Infrastructure

The ancillary facilities and services for the Prosperity Project comprise the following:

· building structures, including the service complex (administration, vehicle maintenance, tire repair, warehouse), assay laboratory and explosive plant;

· power supply from the BC Hydro grid, transmission to site, and project site distribution;

· services, including fresh water supply, fire/fresh water storage and distribution, recycled water collection/storage/distribution, fuel storage and dispensing, sewage collection and treatment, drainage and runoff settling ponds;

· housing facilities for construction personnel and operating employees;

· project site access roads;

· plant site roads, yard areas and parking, and

· security, safety, and first aid facilities

The general site layout site is shown in Figure 18-1.

Figure 18-1 General Site Layout

Site Access Road

Current vehicle access to the site from Williams Lake is via Provincial Highway No. 20 and forestry resource roadways. The road between Williams Lake and the mine site (approximately 180 km) must be an all-weather road and will comprise a portion of the following:

· Provincial Highway No. 20 - 90 km of 2 lane, paved road;

· Taseko Lake (Whitewater) Logging Road - 68 km of one-lane, 5 m wide, gravel road with turnouts;

· 4500 Road (Riverside Haul Road) - 19 km of one-lane, 5 m wide, gravel road with turn-outs;

· and a new Project Site Access Road - 2.8 km of 2 lane, 8 m wide, gravel road.

This road system will serve as the principal access road both during construction and mine operation. Road rehabilitation work will be required on 19 km of the 4500 Road prior to commencement of mine construction. Construction of the 2.8 km Project Site Access Road connection to the 4500 Road will be one of the first activities for the project.

Power Supply and Distribution

B.C. Hydro will supply power to the Prosperity Project via a 125 km, 230 kV overhead transmission line. A Switching Station will be built by B.C. Hydro at Dog Creek adjacent to the existing 230 kV transmission line, which runs south from the Soda Creek substation, near Williams Lake, to the Kelly Lake substation north of Lillooet. The Switching Station will supply the Prosperity Project 230 kV overhead line that will be built and owned by the Prosperity Project. Further engineering details of this 230 kV overhead line are provided in the 2007 report by Ian Hayward International Ltd. and an updated route selection and constraints analysis by HATCH Energy in 2008.

The 230 kV line will be terminated at a structure to be erected within the fenced compound of the main site substation. This will be located close to the concentrator building and will include the main 230 kV circuit breakers protecting three 75/100/125 MVA 230 kV - 25.0 kV three phase oil-filled transformers. The facility will operate on one transformer, with the second unit maintained as an off line spare.

Initially, two generators will be used for construction power. The generators will supply power via a transformer to the 25 kV, metal clad, switchgear located at the Main Substation. Following the mine and process plant commissioning, they will be available to supply emergency power to selected critical loads in the process plant and camp in the event of an outage in the B.C. Hydro power supply.

Power will be distributed throughout the project site via the 25 kV switchgear located in the main substation switchgear building.

An Electrical Load Analysis has been performed based on data extracted from the project equipment list. The analysis utilized motor efficiency and power factor figures at three quarters load from a typical North American manufacturers published data. For all motors, except those of the mills, maximum demand load was assumed to be 90% of the connected load. Thus motors are 10% oversized to allow for actual driven machine load variations over duty cycle, and average demand was set at 85% of the maximum demand load. For the SAG, Ball and Regrind mills, maximum demand will correspond to connected load and average demand will be 85% of the maximum. Standby units are included in the connected load but are assigned a zero demand and load factor.

Mining equipment, especially the shovels, have a power demand curve which varies considerably over the repetitive operating cycle. During the cycle there are periods when regeneration occurs. For this reason, data from a typical manufacturer has been used to represent the mining load as accurately as possible.

In order to calculate annual energy consumption, availabilities of 67% and 92% have been used for the crushing plant and concentrator, respectively. Certain equipment, such as sump pumps, overhead cranes, hoists and other similar service equipment within those plants, can be assumed to operate intermittently.

Lighting loads have been based on building floor areas in conjunction with an estimated load of 3 Watts per square foot. In the buildings, lighting is assumed to be operated on a continuous basis.

Table 18-1 contains the electrical load analysis summary for Year 7.

Table 18-1 Electrical Load Analysis — Year 7

Electrical Summary

		Energy Consumption MWh/y
		based on a 94% motor
Description		efficiency
Water Systems and Water Storage & distribution)		8,187
Fuel Systems		436
Crushing		23,637
Conveyors, Stockpile Reclaim		4,223
Concentrator Building including HVAC		24,560
Grinding		403,147
Pebble Crushing		3,540
Flotation and Air System		71,952
Regrind		53,063
Concentrate Thickening and Loadout		3,807
Concentrator Reagents		1,137
Process Water		5,739
Tailings Pumping		13,348
Water Supply		29,773
Eng $ Admin. Facilities		3,495
Mine Service Facilities		1,461
Assay Laboratory		2,579
Operations Camp		9,867
Plant distribution Losses (5%)		36,218

Total Consumption		700,168

Mill Building and Concentrate Load-out

The grinding circuit, flotation circuits, reagent mixing and concentrate dewatering will be contained in a single building with a total area of approximately 14,000 m2 as shown in Figure 18-2.

A 100/15 tonne crane will service the SAG and ball mill grinding area. Additionally two 20 t cranes will be installed over the flotation/concentrate load-out areas. The large bulk flotation cells will be installed on a slab on grade, inside the mill building. The cleaner flotation cells will be located on elevated steel platforms in the flotation section. Areas with process tanks will be provided with curbs with a spill containment capacity equal to 110% of the contents of the largest tank.

The mill building will be a steel frame structure. The roof will consist of metal sheeting on steel trusses and purlins with an insulated, built-up membrane. The walls will be insulated and metal clad. An interior metal clad liner will be installed on all walls for full height to protect the insulation.

The concentrate storage and load-out areas will follow the same construction methodology and be slab on grade. The concentrate load out area floors will be provided with an under-floor heating system utilizing a pumped hot glycol system to prevent ice build-up on the floor slabs. A front-end loader will load concentrate trucks.

The concentrate thickener and a single stock tank will be located at grade outside the load-out section.

Service Complex

The mobile equipment servicing and maintenance facilities will consist of a service complex and shop, and truck wash. The shop facilities include a small vehicle repair bay, 5 mobile equipment repair bays including lubrication provisions, 1 welding bay, 1 tire shop bay and an electrical bay. The major equipment bays are sized to accommodate 222 t haul trucks and other mine and ancillary mobile service equipment, with 10.5 m wide x 8.5 m high multiplex vertical lift doors. One 50/15 t and two 15 t cranes will service the shop.

The wash bay will be equipped with high-pressure water monitors and steam cleaning equipment. The concrete floor will be sloped towards a drain and an oil interceptor system plus waste oil tank will be included to store residual oils.

The service complex building will be a pre-engineered, steel frame structure, with metal cladding and concrete slab floors. The metal clad roof and walls will be insulated. Interior metal liner will be installed on all walls for part of their height to protect the insulation.

Administration Building and Camp Facility

Taseko will utilize an integrated approach to the design of the administration and camp facility. The administration and change house facilities will be attached to the camp. The administration facility will be contained within a pre-fabricated Sprung structure with the offices located on a mezzanine level and the mine dry, production office and cafeteria located on the ground floor. The camp will be a modular design utilizing container style housing manufactured offsite and delivered to site for final placement.

The materials storage facility consists of two warehouse Sprung structures, one of which will be heated and the other will be for cold storage. These warehouses will be located adjacent to the service complex. A fenced outside storage area will provide 250 m2 of yard storage.

A Sprung structure located at the entrance to the mine site will serve as a fire, ambulance, and Security facility.

The assay and metallurgical laboratory will be located in a separate modular building near the service complex.

Fuel and Lubrication — Delivery, Storage and Dispensing

The diesel fuel storage facility on site will be 300,000 litres providing 3 days of fuel storage when peak consumption of approximately 100,000 liters per day occurs in the middle years of the mine life. The storage volume is considered adequate due to the proximity and the good road transportation network to Williams Lake, where commercial bulk fuel storage facilities are located.

Fire Protection

Fire water protection will be provided for the mill site area and construction/operations camp. During the project construction period, the camp will have an independent fire protection system (storage tanks and pumps) that will be incorporated into the permanent mill site system when it is operational. The primary crusher and overland conveyor will not be provided with fire water protection, because they will be remote from the mill site system and do not have a high fire risk. Fire suppression or retardant system will be provided in the primary crushing building.

Fresh, Process, and Potable Water Supply and Distribution

A suitable fresh water source will be required (non-potable) during construction and operation phases of the project.

The water sources which will be available to the project include:

· Runoff Collection Sump During Operation

· Open Pit Depressurization and Dewatering

· Wells located near the permanent camp

The Process Water Tanks will have a total storage capacity of 110,000 m3 and will be supplied by three sources:

· Runoff Collection Sump Barge Pumps

· Reclaim Water from Tailings Storage Facility

· Reclaim water from the open pit

Steel tanks for fresh, process, and firewater will be situated uphill from the plant site area. Process pumps augment the gravity advantage of the uphill tank positions and will boost the pipeline pressure for distribution and use in the building.

Potable water will be supplied by three proposed wells along the south perimeter of the ultimate open pit or by wells in proximity to the plantsite area. The estimated daily potable water demand during construction will be 200 m3 which is based on a maximum work force of 800 people. During operations, the estimated daily consumption will be 100 m3, which is based on an average onsite work force of 400 people.

Sewage Treatment

Sewage from the mill site and camp areas will be collected by a gravity sewer system, consisting of buried PVC pipes and concrete manholes at all junctions and will be conveyed to a sewage treatment plant. For the concentrator, a sewage lift station and forcemain will be required to pump its sewage to a gravity sewer main. The lift station will be a packaged pump station with a fiberglass chamber and the forcemain will be HDPE pipe.

One sewage treatment plant (STP) will be used to service the mine during the construction phase and continue for operation. The maximum capacity of the plant will be based on a maximum workforce of 800 during construction. Sewage treatment will be by a packaged Rotating Biological Contactor (RBC) unit.

The STP will be located at the west end, low side of the mill site, well away from the camp and other occupied areas.

Communications

Telephone and facsimile communications from the project site will be via microwave to a provincial distribution system. Distribution from the main office at the mine services area will be established across the site via buried lines. Associated equipment will be installed at the camp.

Construction Camp

The construction camp will be located south of the mill site. The construction camp will be constructed in stages in order to accommodate the build-up of workers from the early stage of construction activity to an estimated peak of 800.

On completion of the construction activities, surplus rental bunkhouse units, 1 management complex and portions of the dining and recreation buildings will be dismantled and removed. Any buildings that are retained for the operations camp will be mobilized to the permanent camp and administration facility.

Power supply to the construction camp will be from a 2 - 2200 kW diesel generator set installation. Primary usage of the power will be for heating and lighting.

The propane system will remain in place for the life of the project.

Tailings Containment

Tailings from the concentrator will be collected in a tailings launder and flow by gravity to the tailings containment facility during the first few years of mine operation. A tailings pumpbox and tailings pumps will be required once the level in the containment system has reached an elevation whereby gravity flow will no longer be possible.

Figure 18-3

Tailings Containment

Tailings will be deposited in an impoundment located in the Fish Creek valley upstream from the open pit (Figure 18-3). The Tailings Storage Facility (TSF) has been designed to provide environmentally secure storage for co-disposal of approximately 831 million tonnes of tailings and 858 million tonnes of potentially acid generating (PAG) waste material. The remainder of mine waste generated over the life of the project will be utilized for tailing embankment construction.

The retention earthworks consist of Main, West, and South Embankments. The Main Embankment will be expanded in stages across the Fish Creek Valley, the West Embankment will be constructed along the western ridge which separates the Fish Creek drainage basin from the Big Onion Lake drainage basin, and the South Embankment will be constructed at the southern limit of the TSF, north of Prosperity Lake.

The larger (Main Embankment) containment embankment has a zoned starter water retaining dam. Once the tailings beaches have established a suitable filter, the Main Embankment will be constructed as a free draining structure that utilizes a down stream construction method with a filter and a transition zone supported by the downstream shell zone. This transition is scheduled during Year 3. The Main Embankment is shown in Figure 18-4.

By contrast, the West Embankment will be required by year 3 of operations and will be constructed as a water-retaining structure and raised using the centerline method of construction. Non-PAG waste rock and overburden are planned in the embankment shells as shown in Figure 18-5.

The South Embankment will be constructed as a water-retaining structure and raised using the centerline method of construction starting in year 14.

The discharge of tailings from the delivery pipelines into the TSF will be from a series of large diameter valved offtakes located along the Main and West Embankments. Tailings discharge will begin along the Main Embankment, and will be extended along the West Embankment starting in Year 4. The coarse fraction of the tailings are expected to settle rapidly and will accumulate closer to the discharge points, forming a gentle beach with a slope of about 1 percent.

The material requirements for embankment construction are shown in Table 18-2.

Table 18-2

Embankment Construction Material Requirements

STAGED EMBANKMENT FILL VOLUMES

Print: 12/14/09

	Crest Elev.	Main	West	South	Annual Total
YEAR	(m)	(m3)	(m3)	(m3)	(m3)
-2	1,492	1,582,399	50,000	900,219	2,532,618
-1	1,495	1,018,752	50,000	—	1,068,752
1	1,503	2,480,054	50,000	—	2,530,054
2	1,510	1,876,313	16,453	—	1,892,765
3	1,516	1,977,304	162,129	—	2,139,433
4	1,521	2,291,098	269,860	—	2,560,958
5	1,526	2,105,061	287,794	—	2,392,855
6	1,530	2,383,663	1,065,396	—	3,449,059
7	1,534	2,234,599	1,050,659	—	3,285,258
8	1,538	2,537,182	1,642,468	—	4,179,650
9	1,542	2,191,850	1,418,913	—	3,610,763
10	1,545	2,364,074	935,409	—	3,299,483
11	1,549	2,396,884	1,003,373	—	3,400,257
12	1,552	2,411,476	1,269,313	—	3,680,789
13	1,555	2,178,440	894,432	—	3,072,872
14	1,558	1,868,788	29,015	209,549	2,107,352
15	1,560	1,918,184	41,681	210,603	2,170,468
16	1,563	2,791,276	146,543	274,079	3,211,898
17	1,566	2,300,156	337,085	236,948	2,874,189
18	1,568	2,156,496	959,555	255,144	3,371,195
19	1,570	2,263,532	1,007,181	267,808	3,538,521
20	1,572	2,496,392	1,150,723	337,569	3,984,684
21	1,574	2,482,944	1,149,526	341,040	3,973,510
22	1,576	2,566,588	1,218,711	379,596	4,164,895
23	1,579	2,605,292	1,255,035	401,266	4,261,593
24	1,581	2,567,248	1,248,170	407,572	4,222,990
25	1,582	2,313,384	1,148,408	392,394	3,854,186
26	1,583	1,230,932	611,058	208,789	2,050,779
27	1,584	1,166,176	578,914	197,806	1,942,896
28	1,585	1,159,900	570,140	200,128	1,930,168
29	1,586	1,208,324	569,882	222,877	2,001,083
30	1,587	1,205,656	568,626	222,386	1,996,668
31	1,588	1,204,656	568,154	222,201	1,995,011
32	1,589	1,179,184	556,140	217,502	1,952,826
33	1,590	1,062,800	501,246	196,035	1,760,081
TOTAL		69,777,056	24,381,992	6,301,507	100,460,555

NOTES:

1. THE FILLING SCHEDULE ACCOUNTS FOR 598.3 Mm3 OF TAILINGS AND 438.3 Mm3 OF PAG WASTE ROCK AND OVERBURDEN.

2. 2009 ORE AND WASTE SCHEDULE PROVIDED BY TASEKO MINES LIMITED (DEC 12, 2009).

3. THE DRY DENSITY OF THE TAILINGS IS ASSUMED TO BE 1.2 T/M³ IN THE FIRST 2 YEARS OF PRODUCTION, 1.3 T/M³ IN YEAR 3 AND 1.4 T/M³ TO THE END OF MINE LIFE.

4. PRODUCTION BASED ON 70,000 T/DAY ORE THROUGHPUT (VARIABLE THROUGHPUT IN YR 1 AND 33)

5. AN ADDITIONAL 3M FREEBOARD IS INCLUDED FOR WAVE PROTECTION AND 72 HR STORM EVENT.

6. LEVEL TAILINGS BEACH ASSUMED. PAG WASTE MATERIALS ASSUMED TO BE THE SAME ELEVATION AS THE TAILINGS.

7. PAG WASTE DRY DENSITY IS ASSUMED TO BE 2.04 T/M³ FOR WASTE ROCK AND 1.83 T/M³ FOR OVERBURDEN.

8. VOLUMES ARE BASED ON COMPACTED EMBANKMENT FILL BULK DENSITIES OF 2.3 TONNES/M³.

9. GROUND LEVEL CONSIDERED 1455 M AT THE MAIN EMBANKMENT, 1500 M AT THE WEST EMBANKMENT AND 1530 M AT PROSPERITY LAKE DAM.

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Seepage losses from the Main Embankment will be collected at the Water Collection Pond, which is located between the Waste Storage Area and the Open Pit. Special design provisions to minimize seepage losses include the development of extensive tailings beaches (which isolate the supernatant pond from the embankments), toe drains to reduce seepage gradients and contingency measures for groundwater recovery and recycle.

A Headwater Channel will be constructed in the pre-production years to divert runoff from the undisturbed eastern portion of the Fish Creek catchment area.

Construction of the Stage Ia Main Embankment will start approximately 18 months before mill start-up. Sufficient water will be impounded prior to start-up, and will be available for mill commissioning and early operations. Mill process water for ongoing operations will be reclaimed from the TSF and supplemented with water from the Water Collection Pond.

Details of the site characteristics, geotechnical, hydrogeological and water balance considerations for the tailings facility design, pipeworks, seepage collection and reclamation and closure are contained in the Knight Piésold “Report on Feasibility Design of the 70,000 Tonnes per Day Tailings Storage Facility” dated September 2007.

18.2 Open Pit Design

This section of the report describes the basis for the open pit design including the design parameter basis, design summary, geotechnical considerations, dewatering, and waste material types and storage method.

The open pit design has been based upon the following key considerations:

· Geotechnical recommendations and design criteria, for maximum pit slope and waste dump locations, provided by Knight Piesold Consulting (KP).

· Operating constraints of the equipment selected for mining.

· Minimum mining width defined by shovel double side loading of trucks with allowance for access ramps.

· Bench height achievable and within the safe operating reach of the primary loading unit.

· Minimum haulage road operating width and maximum effective grade within the operating limitations of the primary haulage units.

· Logical and efficient scheduling of material movement from multiple phases of pit expansion to the crusher, the stockpiles and to final waste material placement sites.

Geotechnical investigations and testing were undertaken by KP. The complete test results, findings and recommendations for the pit wall slopes, waste dumps and results of hydrological investigations are contained in the KP report, “2007 Feasibility Pit Slope Design”, dated September 21, 2007. KP’s work consisted of site reconnaissance and mapping, oriented core

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diamond drilling and detailed logging of fracture data, in-situ permeability testing, point load testing, uniaxial compressive and tri-axial strength tests and direct shear tests on rock joints.

Geotechnical core logging data were used to develop a rock mass classification system and rock mass model for the deposit. Mapping data were used to determine structural discontinuities and to assess the potential for wedge and plane failures in the pit walls. These assessments were the basis for stability analyses of failure modes along structural discontinuities and for evaluation of deep-seated failure.

The mine design drawings and ore reserve reporting was completed by the engineering staff of Taseko Mines Limited and AKF Mining Services Inc. Mining phases were smoothed, and haulage roads, stockpiles and waste dumps were located.

Geotechnical Considerations

The Prosperity open pit will be nominally 800m deep when complete. Open pit wall slope stability is dependent upon the following site specific factors:

· Geological structure

· Rock alteration

· Intact rock strength

· Rock stress

· Groundwater conditions

· Discontinuity strength and orientation

· Pit geometry

· Blasting practices

· Climatic conditions

· Time

In general the rock mass quality at Prosperity ranges from fair to good. There are two major faults within the pit limit. These are referred to as the QD and East Faults. These structures are near vertical, sub-parallel and trend North-South through the center of the deposit. There do not appear to be any major structures that will adversely influence the stability of the pit slopes.

The Prosperity Deposit is centered about a diorite intrusive where potassic alteration is associated with the core of the mineralized zone. This central zone of mineralization is surrounded by a propylitic alteration zone. A retro-grade phyllic alteration is overprinted on the propylitic and potassic zones. Within the potassic zone there is a well defined vertical zonation defined by dissolution of gypsum on joint surfaces. The “gypsum” line defines the change from generally competent rock to competent rock and is used to separate structural domains for the purposes of mine design.

Intact rock strength is an important consideration, as many potential failure surfaces are not completely developed and require some failure of intact rock. The moderate to high strength of the rock at Prosperity site is beneficial due to the high stresses that are expected to develop

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in the pit slopes during later stages of mining. The uniaxial compressive strength, based on point load tests, varies but averages 112 megaPascals (mPa).

The rock stress conditions within the rock mass are a significant factor for high slopes. KP has used a sophisticated finite difference computer model (FLAC) to assess the potential overstressing of the rock in the proposed pit slopes.

The predominant jointing patterns are sub-vertical and coincident with the main vein systems. Secondary veins have also been identified dipping out of the East pit slopes. KP has investigated the potential for adversely oriented structural features at depth at or near the final pit walls. The finding of this investigation was that there is a very low likelihood of adverse structures in the form of open joints. Structural features in close proximity to final walls will be primarily quartz and sulphide veins.

Based upon three structural domains the open pit has been divided by KP vertically into three major slope design sectors that correspond with:

· Sector I - Surface materials including overburden and basalt

· Sector II - Upper Zone located above the “gypsum line”

· Sector III - Lower Zone located below the “gypsum line”

These major sectors have been further subdivided in detail as shown in Figure 18-6. However, the actual design recommendations for each major sector are for the most part identical and are summarized in Table 18-3. The overburden will be mined leaving a 30° inter-ramp slope. The basalt formation on surface will be mined leaving a 45° inter-ramp slope. The Middle Zone will primarily be mined leaving a 45° inter-ramp slope and the Lower Zone inter-ramp slope will be increased to 50°.

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Figure 18-6

Geotechnical Pit Slope Design Sectors Plan

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Table 18-3

Recommended Wall Slopes

Design for Near Surface Materials

Inter-Ramp Slope

Bench Height

Berm Interval

Berm Width

Interberm Slope

Design Sector

Geologic Domain

degrees

Overburden

30.0

15.0

8.0

40.0

Overburden-Basalt

45.0

15.0

30.0

8.0

65.0

Design Above the Gypsum Line

		Inter-Ramp Slope	Bench Height	Berm Interval	Berm Width	Interberm Slope
Design Sector	Geologic Domain	degrees	m	m	m	degrees
IIa	Middle West Sector	45.0	15.0	15.0	8.0	65.0
IIb	Middle West Sector - Potassic	45.0	15.0	15.0	8.0	65.0
IIc	Middle Northwest Sector	45.0	15.0	15.0	8.0	65.0
IId	Middle North Sector	30.0	15.0	15.0	8.0	65.0
IIe	MiddleNortheast Sector	30.0	15.0	15.0	8.0	65.0
IIf	Middle East Sector	45.0	15.0	15.0	8.0	65.0
IIg	Middle Southeast Sector	45.0	15.0	15.0	8.0	65.0
IIh	MiddleSouth Sector	45.0	15.0	15.0	8.0	65.0
IIi	Middle Southwest Sector	45.0	15.0	15.0	8.0	65.0

Design Below the Gypsum Line

		Inter-Ramp Slope	Bench Height	Berm Interval	Berm Width	Interberm Slope
Design Sector	Geologic Domain	degrees	m	m	m	degrees
IIIa	Lower West Sector	45.0	15.0	15.0	8.0	65.0
IIIb	Lower Northwest Sector	50.0	15.0	30.0	11.0	65.0
IIIc	Lower North Sector	50.0	15.0	30.0	11.0	65.0
IIId	Lower Northeast Sector	45.0	15.0	15.0	8.0	65.0
IIIe	Lower Southwest Sector	50.0	15.0	30.0	11.0	65.0
IIIf	Lower South Sector	50.0	15.0	30.0	11.0	65.0
IIIg	Lower Southwest Sector	50.0	15.0	30.0	11.0	65.0

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Slope stability analyses using a RMR rock mass classification basis for limit equilibrium and numerical modeling have been undertaken as part of the process of defining the design slope recommendations. These analyses have indicated a potential for instability in the zone of potassic alteration located in the central west slope. The analyses also indicated that stability was sensitive to groundwater depressurization and draw down. As a result increased drain hole density and hole lengths will be required in this area.

The climatic conditions at the Prosperity Project are typical of the British Columbia Chilcotin District with an annual average of 524 mm of rain equivalent precipitation. The seasons in this area are well defined with relatively predictable periods of “freeze up” in the fall and “break up” in the spring. The “break up” period is characterized by increased water flow from melting snow and cyclical thawing and freezing of the surface materials on pit slopes. This action results in decreased slope stability particularly at the smaller bench scale where there will be a marked increase in small face failures and raveling of rock.

The ultimate pit geometry is roughly oval and the internal pit phases expand in all directions about the Phase 1 Starter Pit. As such, during the life of the mine all internal walls are temporary and will be mined. Final walls will occur only in the Phase 5 Pit that is active for a period of 25 years between Year 6 and Year 31 of the production schedule. Phase 1 and Phase 2 pit walls will typically be exposed for 2 years and the Phase 3 walls will be exposed for 4 years prior to excavation.

Drilling and blasting near both temporary and final walls will require buffer blasting. KP have recommended overall wall slopes of 30° in overburden, 45° above the “gypsum line” and 50° below the “gypsum line”. The KP recommendations for bench and berm configuration were based upon single benching and achieving steep inter-berm face angles up to 75°. The designs incorporated in this study assume that double benching will be possible and that shallower inter-berm angles to 64° will be allowed resulting in berm widths from 10 to 15 m width. This assumption simplifies the wall control blasting requirement and the necessity for multiple hole sizes and drill rig configurations on wall control blasts.

Buffer rows will be drilled using production blasthole drills at a reduced spacing and with adjustment to hole depth near design berm crests. Line holes may be incorporated in the blast patterns. Walls will be scaled carefully on each bench.

Pit Dewatering

Mine dewatering has been addressed by KP and is summarized in this section of the report in terms of how it relates to the mine operations. The water management recommendations are the basis for capital and operating cost estimates in later sections of the report.

It has been recognized that the open pit development will have significant impact on the local hydrogeological regime, as the pit will become a groundwater discharge area. The water table is currently at or near the ground surface and provisions have been made for an extensive slope depressurization system. Groundwater dewatering wells and slope depressurization will be concentrated in Sector IIIc and Sector IIIf shown in Figure 18-6.

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Pit inflows will likely be dominated by localized confined aquifers in the southern area of the pit from zones of higher rock mass permeability related to major structures and from unconfined flow in the upper 150 to 300 m of fractured rock mass above the gypsum line. Inflows from good quality, low permeability rock below and peripheral to the gypsum line are expected to be low.

Depressurization systems are important in overall pit slope design. A combination of techniques including vertical wells, in-pit horizontal drains and collection systems will be implemented as a staged approach during pit development.

The open pit dewatering system has been designed to meet the combined requirements of the expected groundwater pit inflow rates and runoff from precipitation. The annual contribution of direct precipitation to the in-pit pumping requirements has been estimated for the average annual precipitation volume with a ten year return period, and the storm flow rate required to remove ponded water from the one in ten year, 24 hour storm event within 96 hours. The peak operational design capacity of the system is 400 litres/second.

General Design

The mining equipment will operate on a 15m high bench in overburden and hardrock. Wall slope design changes will be implemented by varying the berm widths and inter-berm slope angles.

Berms will be left on every bench in overburden and on alternate benches in hardrock. Berm width design will vary from 15 m to 10 m as the overall wall slope is increased from 45° to 50° in the Lower Zone. The general mine design parameters are summarized in Table 18-4.

The open pit will be mined in five phases commencing with the Phase 1 Starter Pit. The pit will be partially pre-stripped during the preproduction development period. The Starter Pit will provide building materials for the tailings impoundment starter dam. The Phase 2 through Phase 5 pits are radial expansions of the mine about the Starter Pit creating a progressively deeper pit.

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Table 18-4 Design Parameters

Open Pit Design

Bench and Berm Design
Bench Height	metres	15.0
Bench Interval Overburden	metres	15.0
Bench Interval Gypsum Zone	metres	15.0
Bench Interval Hardrock	metres	30.0

Interbench Face Angle Design Sector I & II	degrees	65°
Interbench Face Angle Design Sector III	degrees	65°

Haulroad Design
Total Road Allowance	metres	30.0
Maximum Haulroad Grade	percent	10.0%

Minimum Pushback Width
Minimum Pushback	metres	80.0

Waste Dumps & Stockpiles

Material Properties
Overburden Bulk Density Placed	t/pcm	1.83
Waste Rock Bulk Density Placed	t/pcm	2.04

Seismic Criteria	year	1 in 1000
Maximum Design Earthquake	g	0.1
Acceleration

Pile Stability Criteria
Minimum Factor of Safety During Operations		1.2
Minimum Factor of Safety for Closure		1.5
Minimum Factor of Safety for Seismic Loading		1.0

Final Slopes
Overburden Bench Height		30.0
Overburden Berm Width		20.0
Overburden Face Slope		1.3:1
Waste Rock Bench Height		30.0
Waste Rock Berm Width		20.0
Waste Rock Face Slope		1.2:1

t/pcm = tonnes per placed cubic metre

The minimum pushback width is 80 m; however in general the expansions are in excess of 100 m width. Haul road allowances have provided at 35 m. Roads are designed at a maximum of 10% grade and are located to spiral counterclockwise into the pit bottom.

The ultimate pit features are summarized as follows:

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· 2,130 m E-W by 1825 m N-S

· Final ramp exit elevation 1475 m

· Ultimate pit bottom elevation 765 m

· Maximum wall height — 800 m in the SW quadrant with maximum elevation 1560 m

· Final overall wall slope angles in the following directions:

	North Wall	43°
	East Wall	43°
	South Wall	42°
	West Wall	42°

Waste Storage, Stockpiles, and Roads

The waste or non-ore material types included in the reserves and material movement schedules are subdivided into overburden, waste rock and stockpiled ore. The waste materials are further subdivided into potential acid generating (PAG) and non-potentially acid generating (non-PAG) proportions. Classification of waste materials is based on spatial remodeling of the ABA data by Taseko in 2009. Sub-aqueous storage is proposed for PAG overburden and waste rock and the balance of the overburden and waste can be used for construction purposes or placed on surface storage sites where surface drainage is controlled and treatable as required.

A total of 2.0 billion tonnes of material will be mined from the open pit, including 771 million tonnes of ore directly to the crusher, 60.0 million tonnes of stockpiled ore, 25 million tonnes of PAG overburden, 201 million tonnes of non-PAG overburden, 832 million tonnes of PAG waste rock and 157 million tonnes of non-PAG waste rock.

The area underlying the overburden and waste dump site is characterized by up to 20 m of glacial till, which overlies Quaternary Glaciofluvial and Glaciolacustrine units. These in turn overlie Miocene Basalt flows and a Miocene Glaciofluvial Unit followed by glacial till and colluvium. These units extend south and overlie the open pit area as well.

Overburden has been classified as PAG and non-PAG in nature. The PAG overburden contains weathered rock which includes oxidized or partially weathered sulphide minerals. This material will be placed in the tailings management facility. Non-PAG overburden will be placed in the overburden stockpile located to the south of the open pit.

Overburden piles will be developed in 30m high benches, each offset from the downstream edge by 20 m to provide an overall slope of approximately 2H:1V (2 m height to 1 m vertical). Each bench will be constructed from lifts of approximately 15 m by end dumping the material to form an angle of repose slope angle of approximately 1.3H:1V.

The total PAG waste is 858 million tonnes. This quantity of material will be hauled from the mine and placed in the tailings management facility. It will be dumped in lifts and dozed out into the area of active tailings deposition.

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The total non-PAG waste is 358 million tonnes. Over the life of the mine 231 million tonnes will be required for construction in specific zones of the tailings dams. The filling of Fish Lake to the 1470 elevation will require approximately 17 million tonnes. This will be progressively be filled in with layers of durable free draining waste rock.

The total ore quantity stockpiled in the variable cutoff grade production schedule is a total of 60 million tonnes.

As described above, the dumps and stockpiles will be constructed in lifts with berms left at 30 m intervals. Overall final slopes will be 2H:1V and crests will be contoured for reclamation. Prior to placement of overburden and waste in the stockpile areas the vegetation will be cleared, and diversion & runoff collection ditches will be constructed.

Stability analyses have been carried out for waste rock and overburden piles. The analyses were performed for static and seismic conditions. A limit equilibrium method of calculation was used. Minimum factors of safety for the static state were 1.5 for both overburden and waste rock and 1.2 for seismic conditions.

Haul roads will be required from the mine to the crusher, stockpiles, overburden spoil piles, waste dumps and the tailings management facility for construction and waste disposal. These roads will be constructed with materials derived from mine operations. The mine haul roads will be designed on a 35 m road allowance with a running surface of 3 times the total operating width of the haulage truck. The haulage roads will not exceed a design grade of 10%. They will be built with an operating surface of 30 m and additional allowance for ditches and berms where required.

18.3 Mining Operations

Production Schedule

The geometry of the ore deposit in relation to the relatively gentle topographic surface allows for flexibility in pit sequencing. Ore is easily accessible near surface in the Phase 1 Starter Pit. Once the relatively thin layer of overburden is removed the ore is released in the Phase 1 Pit. Subsequent pit expansions are generally symmetrical about the Starter Pit and as such become progressively larger and deeper. The sequencing therefore becomes an exercise of balancing total production over the life of the mine to defer major stripping and associated capital equipment while effectively utilizing equipment on site.

The proposed final pit is shown in Figure 18-7.

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Figure 18-7

Prosperity Final Pit — End of Year 31

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The mine production forecast has been derived by scheduling ore at a variable declining NSR cutoff.

Ore has been scheduled to provide 25.5 million tonnes of ore to the primary crusher annually. The mine will operate 365 days per year with a nominal crusher throughput of 70,000 tonnes/day. The production schedule is shown in Table 18-5 Mine Production Forecast. This data is represented graphically in Figure 18-8 Material Movement Schedule.

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Table 18-5 Mine Production Forecast


	Year	-2		-1		1		2		3		4		5		6		7		8		9		10		11		12		13		14		15		16		17
Cutoff ($NSR/t)		$	0.00	$	0.00	$	11.50	$	11.50	$	11.50	$	9.00	$	9.00	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50
Ore to Mill	(t x 1000)			—		15,318		25,567		25,564		25,565		25,558		25,570		25,590		25,560		25,549		25,606		25,607		25,580		25,544		25,533		25,594		25,593		25,560
Gold Grade	g/t			—		0.47		0.51		0.55		0.47		0.51		0.42		0.38		0.38		0.34		0.34		0.34		0.37		0.37		0.40		0.43		0.48		0.46
Copper Grade	%			—		0.24		0.25		0.26		0.23		0.24		0.19		0.18		0.19		0.17		0.17		0.17		0.19		0.20		0.22		0.23		0.25		0.24
Ore to Stockpile	(t x 1000)			165		7,389		18,214		15,350		8,231		10,584		—		—		—		—		—		—		—		—		—		—		—		—
Gold Grade	g/t			0.29		0.32		0.31		0.31		0.28		0.27		—		—		—		—		—		—		—		—		—		—		—		—
Copper Grade	%			0.19		0.18		0.17		0.16		0.14		0.14		—		—		—		—		—		—		—		—		—		—		—		—
Total Waste	(t x 1000)	6,000		14,000		52,302		31,276		34,153		41,195		38,912		49,461		49,524		49,336		49,456		49,322		49,335		49,238		49,326		49,204		49,382		49,421		49,266
Total Material Moved	(t x 1000)	6,000		14,165		75,009		75,057		75,067		74,991		75,053		75,031		75,114		74,895		75,004		74,928		74,942		74,818		74,870		74,737		74,976		75,015		74,827


	Year	18		19		20		21		22		23		24		25		26		27		28		29		30		31			32			33			34	Total
Cutoff ($NSR/t)		$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50	$	5.50		$	5.50		$	5.50
Ore to Mill	(t x 1000)	25,529.0		25,552.9		25,599.1		25,527.2		25,529.9		25,528.9		25,523.5		25,616.8		25,651.9		25,598.1		25,534.4		25,543.1		25,570.9		14,520.6			—			—			—	771,286
Gold Grade	g/t	0.37		0.37		0.37		0.37		0.37		0.37		0.39		0.42		0.43		0.44		0.48		0.50		0.53		0.52			—			—			—	0.42
Copper Grade	%	0.21		0.21		0.21		0.21		0.21		0.22		0.23		0.25		0.26		0.28		0.30		0.32		0.35		0.37			—			—			—	0.23
Ore to Stockpile	(t x 1000)	—		—		—		—		—		—		—		—		—		—		—		—		—		(11,029		)	(25,550		)	(23,353		)	—	59,932
Gold Grade	g/t	—		—		—		—		—		—		—		—		—		—		—		—		—		0.16			0.16			0.16			—	0.30
Copper Grade	%	—		—		—		—		—		—		—		—		—		—		—		—		—		0.30			0.30			0.30			—	0.16
Total Waste	(t x 1000)	49,379		49,454		49,193		49,268		49,555		49,500		49,465		41,842		8,033		7,013		1,247		921		712		161
Total Material Moved	(t x 1000)	74,908		75,007		74,793		74,795		75,085		75,029		74,988		67,404		33,630		32,556		26,782		26,464		26,283		25,711			25,550			23,353				2,106,838

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Figure 18-8 Material Movement Schedule

Year 1 ore production has been reduced to 15.3 million tonnes to allow for startup and commissioning. Annual waste and overburden quantities have been calculated according to the strip ratio of the scheduled benches. A total of 771 million tonnes of ore will be mined and hauled directly to the primary crusher.

Ore production from the open pit will cease in Year 31 of the current mine plan. Recovery of ore from stockpile will sustain mill production through year 33. A total of 60 million tonnes will be recovered from stock pile at the end of the pit operating life.

Equipment Selection

The mine equipment has been selected given the following considerations:

· The simultaneous distribution of multiple operating faces at several locations determined by the long range production schedule

· The necessity to minimize unit operating costs by using large scale mining equipment

· Use of well proven equipment technology and coordination of operating machines using advanced systems

· Use of equipment assembled with modular components in order to minimize onsite maintenance allowing maintenance personnel to focus on servicing and component replacements.

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The mine will operate using electric cable shovels and rotary drills. Diesel electric trucks and a support equipment fleet will gradually be increased to match the production schedule that will average total production in Years 1 through 24 at 200,000 tpd.

In general it is expected that the major equipment will have an effective operating time of 85% corresponding with a 51 minute hour. The available versus scheduled time will be 10.2 hours available per shift or 85% of the shift. Detailed equipment productivity calculations have been made on an annual basis for drills, shovels and trucks. Support equipment operating time has been factored on an annual basis according to the total annual material movement.

Mechanical availability of the individual drills, shovels and trucks were fixed over the life of the mine. Operating costs, discussed in Section 18 have been calculated on a life cycle basis. The major mining machines including blasthole drills, shovels and trucks are expected to be replaced during the life of the mine.

The primary blasthole drills will be electric powered rotary machines capable of drilling 311 mm holes.

The loading fleet will consist of a total of 3- 43.0 m3 capacity electric cable shovels. Initially, a loader will be required in preproduction, prior to availability of electric power, with shovels coming into production in Year 1. The loader will then be available to work in stockpile areas, low face conditions and where required to meet production objectives during periods of unscheduled shovel downtime.

The haulage trucks selected are 222 tonne capacity diesel electric off road end dump units. The selected truck fleet matches the loading units and overall haul profiles. Truck additions will be made as required in each year until the fleet total of 35 trucks is reached in Year 18 of the mine plan.

The mining support equipment includes track dozers, wheel dozers, graders, water trucks, and scrapers required for road, bench and dump maintenance. Miscellaneous ancillary equipment is also required to service, maintain the major equipment and support ongoing pit operations.

Explosives will be delivered to the blasthole by a contracted supplier. The blasting crew will require support equipment to pump wet holes, deliver blasting accessories and stem holes. The bulk delivery trucks and storage facilities will be provided by the explosives contractor.

Stockpiled Ore

Ore which is not taken to the mill for processing during the early years of the project will be hauled by truck to a storage area close to the primary crusher. This stockpile will be re-handled and processed when the pit has been mined out or during times of low pit ore release.

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Mined Waste Rock

Mined waste rock will be classified as either PAG or non-PAG based on the potential for acid generation.

PAG waste rock will be hauled from the mine and placed in the tailings storage facility in order to prevent oxidation by isolating it from the atmosphere through submersing.

Non-PAG waste rock will be used as construction materials in roads, dam embankments, and platforms for the low grade and overburden stockpiles.

18.4 Processing and Concentrator

The process design criteria have been developed based on extensive metallurgical test work and current industry operating experience in the processing of copper-gold ores. Further, as a direct result of the ability of the mine to support a large production rate, the process design criteria have incorporated the latest industry proven technology for large scale comminution and recovery equipment. The benefits of utilizing large capacity process equipment will be realized in lower process operating unit costs.

Specific aspects of the design criteria are:

· Single large, industry proven primary semi-autogenous grinding mill

· Two large, industry proven secondary ball grinding mills.

· Seven vertical stirred regrind mills

· Large scale flotation cells

· Bulk handling systems for concentrate

· Process automation

The process design is conventional and consists of SAG and ball mill grinding; bulk sulphide flotation, regrind and bulk rougher/scavenger cleaner flotation, cleaner flotation and concentrate dewatering. A simplified flowsheet is shown in Figure 18-9. The concentrator is designed to operate 24 hours per day, 365 days per year.

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The primary SAG mill size was based on MacPherson AG (Autogenous Grinding) work indices derived from test work. The ball mill sizes were based on standard Bond Ball Mill work indices also determined from test work. An evaluation of the large scale processing equipment available with respect to the test work requirements indicated that a nominal processing capacity of 70,000 tonnes/day could be achieved.

The bulk of the metallurgical test work was performed at Lakefield Research under the supervision of Melis Engineering. Semi-autogenous grinding amenability test work using the MacPherson test procedure was performed by Hazen Research. Bond ball mill grinding test work was performed by Lakefield. Flotation test work performed by Lakefield consisted of batch tests, locked cycle tests and pilot plant runs. The sample material used for the tests was either half-core or assay reject samples from the upper, middle and lower zones of the ore body.

Based on three half-core tests, a cost benefit analysis using the value of incremental copper and gold recoveries against net grinding power costs concluded that the optimum 80% cumulative passing size (K80) of the primary grind was approximately 160 µm.

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Five batch tests were completed to analyze the effects of regrind particle size distribution on copper and gold recoveries to the third cleaner concentrate. Based on this test work, Melis recommended that the optimum regrind K80 was between 14 and 17 µm.

Historical analysis of the regrind test work results suggests the copper and gold recoveries may actually improve when the regrind becomes slightly coarser. Test work performed by G & T Laboratories in November and December 1998 on Prosperity concentrate also indicated that a coarser grind may achieve increased recoveries, especially with respect to gold. In light of this data and taking into account regrind ball mill size considerations, a regrind K80 of 19 µm had been specified in the design criteria. Additional grind-recovery test work was conducted in 2008 to confirm the optimum regrind K80. Varying grind sizes and alternate regrind flow sheets were evaluated at Process Research Associates in Richmond, British Columbia on core from the 2007 drilling campaign that represented years 1 through 4 mill feed. This test work did not result in improved metallurgical performance, and a 17 um regrind design target was carried forward. This plant design can readily be operated at regrind values coarser than the 17 um regrind design target if desired.

The copper and gold head grades in the design criteria reflect those presented in the life of mine production schedule (see Section 18.2). The projected recoveries presented in the design criteria have been determined by an evaluation of the test work results and the mine head grades. Based strictly on the test work results, Melis estimated a target recovery for copper of over 90% for all three ore zones. Analysis of the flotation test work results generally concur with the Melis test work. However, for actual plant operation projected average copper recoveries are 87.0, 88.7 and 90.2 for upper, middle and lower zones respectively, though projected recoveries will vary depending on head grade and mineralogy. Plant recoveries are approximately 2 percent lower than the test work recoveries and reflect industry experience where flotation plant recoveries tend to be lower than laboratory scale test recoveries. These lower plant recoveries are a result of process inefficiencies realized in actual plant operating conditions.

In flotation test work the gold recoveries showed significant variation which has made it difficult to predict the expected recovery. This difficulty is compounded by the lack of comparable tests performed under similar conditions at Lakefield. However, analysis of the test work indicated that the gold recovery projections presented by Melis may be conservative and these recoveries have been used in the design criteria. Additional test work could be performed to more accurately predict the recovery of gold in the flotation circuit, but given the conservative nature of the prediction is not required.

Predicted recoveries and grades as shown in Table 18-6 were used as a basis for the process and concentrator design.

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Table 18-6

Predicted Recoveries & Grades

			Concentrate
	Head Grade		Grade		Recovery %
Zone	% Cu	gpt Au	% Cu	g/t Au	Cu	Au
Upper	0.217	0.485	24.31	44.7	87.0	71.6
Middle	0.233	0.441	24.6	40.4	88.7	77.0
Lower	0.310	0.496	25.8	35.9	90.2	78.4

18.5 Processing, Crushing and Ore Reclaim

The Prosperity crushing circuit has been designed to crush on average 70,000 t/d of run-of-mine ore from minus 1,000 mm to 100% passing 350 mm and 80% passing 150 mm at a nominal rate of 4,444 tonnes/hr. The crushing system will consist of a single gyratory crusher operating at 80% utilization per 21 hour day, 365 days per year basis. The primary crusher will be located near the east corner of the ultimate pit boundary.

Product from the crusher will discharge into the surge pocket directly below. Crushed material will be drawn from the surge pocket by a variable speed apron feeder which will discharge onto an overland conveyor which will transport the crushed ore from the primary crusher approximately 2,000 m to the coarse ore stockpile. The coarse ore stockpile is estimated to have a live capacity of 55,000 tonnes. A belt weigh scale, and a tramp iron magnet to remove tramp metal, will be installed on the conveyor.

The primary crusher will have a 10 tonne auxiliary hoist for maintenance purposes. A hydraulic rock breaker will be provided at the crusher dump pocket for breaking any oversize run-of-mine material.

Crushed ore will be reclaimed from the coarse ore stockpile via three variable speed apron reclaim feeders located in a concrete reclaim tunnel. The three apron feeders will feed directly onto the SAG mill feed conveyor.

Grinding

The primary grinding circuit will be designed to process an average of 70,000 tonnes/day at 92% utilization on a 24 hour per day, 365 day per year basis. The grinding circuit will be designed to reduce the crushed ore from 80% passing 150 mm to 80% passing 170 µm. The grinding operation will consist of a SAG mill and two ball mills.

The SAG mill will be 12.2 m in diameter (inside shell) by 6.1 m long (effective grinding length) and will be driven by a 21,000 kW variable speed wrap-around gearless drive motor. The SAG

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mill has been designed to operate with a nominal ball charge of 11 percent by volume. The product from the SAG mill circuit is projected to be 80% passing 2,000 µm. The SAG mill will discharge via a screen into the cyclone feed pump box. This cyclone feed pump box receives material from the SAG and Ball Mills, and feeds the two cyclone clusters. The screen oversize is returned to the SAG mill conveyor by a series of three conventional conveyors. Provision is made to install a pebble crusher in the future if required.

Each of the two ball mill circuits consists of a 7900 mm diameter by 14600 mm long ball mill with dual pinion drive. The motors are 18,000 kW wrap around motors.

The ball mills each discharge into the cyclone feed pump box and individual variable speed pumps feed a dedicated cluster of 12 cyclones for each Ball Mill. The cyclone underflow returns to the ball mill while the cyclone overflow discharges to the flotation circuit.

The 80% passing 200 µm cyclone overflow will flow by gravity to the bulk flotation circuit. The cyclone underflow will be returned as ball mill feed. An on-line particle size indicator (PSI) is proposed for monitoring of the cyclone overflow streams and to provide grinding process control.

A weigh scale will be provided on the SAG mill feed conveyor for controlling, monitoring and recording the concentrator fresh feed rate. A weigh scale will be provided on the SAG mill screen oversize conveyor to monitor SAG mill circulating loads.

Bulk Rougher/Scavenger Flotation

The bulk flotation circuit will consist of two parallel trains of bulk rougher cells.

Cyclone overflow from each of the two parallel grinding circuits will flow by gravity into a crossover box and then split into two equal portions to feed each of the two rows of seven cells. The cells in each row are proposed to be approximately 300 m3 forced air cells, giving a total of 14 cells for the complete bulk rougher flotation area. This combination of cells provides the power and retention time required for flotation. The tailings from the last scavenger cells in each row are combined as bulk flotation tailings.

The flotation reagent distribution system will permit the addition of flotation reagents, collector, promoter and frother at each cell in the bulk sulphide flotation circuits. Flotation air for all flotation sections will be provided by air blowers with contingent capacity for a standby blower to allow for operational main, that will supply air to all flotation cells via a ring-main system. The air supply for each cell will be controlled by valves located at specific locations.

On-stream analyzers will be installed to provide on line analysis of the flotation products. This data will be monitored continuously and used to provide optimum process control for the flotation circuits.

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Regrind and Bulk Cleaner Flotation

The concentrate from the bulk rougher/scavenger flotation circuit is combined with the bulk cleaner scavenger concentrate as feed to the regrind circuit. In order to provide the liberation required to achieve the design copper concentrate grade and copper and gold recovery in the concentrate, the regrind circuit has been designed using a target P80 of 17 to 20 µm. The regrind circuit is designed with seven parallel vertical stirred regrind mills. Each regrind mill circuit consists of a 1,120 kW stirred mill operating in closed circuit with a cluster of seven 250 mm cyclones. The cyclone overflow at 20% solids flows to the bulk cleaner flotation circuit.

The bulk cleaner circuit consists of a single row of four approximately 160 m3 forced air flotation cells followed by a further three approximately 160 m3 forced air bulk cleaner scavenger flotation cells. The bulk cleaner concentrate is pumped to the copper first cleaner flotation circuit while the bulk cleaner scavenger concentrate is returned to the regrind circuit.

The bulk cleaner scavenger tails are combined with the bulk rougher scavenger tailings to form the final plant tailings.

Cleaner Flotation

The bulk cleaner concentrate is cleaned in three stages of copper cleaning at an elevated pH to produce the final copper concentrate. The tailings from the copper cleaning circuit are returned to the head end of the bulk cleaner flotation circuit.

The copper first cleaner flotation uses ten approximately 20 m3 forced air flotation cells. The concentrate is pumped to seven approximately 20 m3 forced air flotation cells. The copper first cleaner tails are pumped to the bulk cleaner circuit. The copper second cleaner concentrate is pumped to the copper third cleaners and the tailings are returned by gravity to the copper first cleaners.

The copper third cleaners consist of three approximately 20 m3 forced air flotation cells. The third copper cleaner concentrate is the final copper concentrate, which is pumped to the concentrate thickener, and the copper third cleaner tailings are returned by gravity to the copper second cleaners.

Concentrate Dewatering and Loadout

Final copper-gold concentrate from the cleaner flotation circuit will be pumped to the concentrate dewatering and load-out circuit. The concentrate dewatering circuit will consist of thickening and filtering unit processes.

Concentrate from the cleaner flotation circuit will be pumped to the 16 m diameter high capacity concentrate thickener. The concentrate will be thickened to approximately 60% solids by weight and pumped to one 5.0 m diameter x 10.0 m high agitated concentrate stock tanks that will provide approximately 5 hours of total storage capacity. Overflow from the concentrate thickener

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will flow to a standpipe and will be pumped to the distribution points in the milling and flotation areas.

Concentrate from the stock tanks will be pumped to an approximately 120m2 automatic pressure filter where the moisture content of the concentrate will be reduced to an estimated 8%. Filtrate from the concentrate filter will be combined with the thickener feed and gravity flow to the concentrate thickener feed box.

Concentrate filter cake from the pressure filter will discharge (free-fall) directly onto stockpiles located on floor level below the filters. The concentrate will be reclaimed by a front end loader and loaded directly into bulk concentrate highway transport “B” trailers. The concentrate trailers will report to a truck wash station to remove and recover concentrate residue and spillage prior to exiting the load-out facility. This measure is intended to minimize potential environmental concerns. The trucks will report to a weigh scale to determine the concentrate load and will transport the concentrate to the Gibraltar concentrate load-out facility for transfer to the CN rail transport system.

Process, Fresh, and Fire Protection Water

The concentrator water system will consist of four water distribution systems: process, fresh, potable and fire protection.

The Prosperity water supply system has been designed to maximize the recycle of process and mine waters for re-use in the process. Process water will be made up of water recovered from the concentrate dewatering system, tailings containment pond, run-off water pond and pit depressurization wells. All process water will be stored in the process water tank prior to being gravity fed to a header inside the concentrator building. Process water will provide the bulk of water required for process and clean-up purposes. The process water will be pumped from the header through the concentrator in a ring-main for distribution to all points of use.

Potable water will be supplied from 3 deep aquifer wells situated above the plant site. The water will be pumped to the potable water tank. Two distribution pumps (1 operating, 1 stand-by) will deliver water to the primary crusher, concentrator and the camp.

Fresh water will be supplied from 12 shallow and 9 deep pit depressurization wells. Fresh/raw water will be stored in the fresh water tank. Fresh water pumps (1 operating, 1 stand-by) will supply water to the dust collector extraction fans and mill heat exchanger. Gland water pumps (1 operating, one stand-by) will supply water to the gland seals of the process slurry pumps.

Fire water will be stored in a separate fire water tank. Water will be reserved for fire water, which will provide the required 1 hour of reserve at 340 m3 per hour consumption. Fire water will be pumped through the concentrator fire water ring-main and into the fire hydrant perimeter ring-main. An electric jockey pump will maintain fire water pressure in the fire sprinkler system. An electric fire water pump will be provided to supply the fire water distribution system. A diesel powered fire water pump will be installed to provide back up pumping capability in the event of loss of electrical power.

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Reagents and Services

Reagent mixing, storage and distribution systems will be provided for the lime, primary collector, secondary collector, frother, flocculant, depressant, and spare reagent.

Quick lime will be delivered in bulk to the mill site in self unloading trucks and will be offloaded into a lime silo via a pneumatic transfer system. The lime will be fed from the silo via a screw feeder into a lime slaking ball mill. Lime consumption is projected to be in the range of 530 gpt.

The proposed flotation collectors, based on the metallurgical test program are xanthate based as the primary collector and thionocarbamate and/or dithiophosphate based as the secondary collector.

The secondary collector will be pumped from the bulk container truck into the 4.50 m diameter x 4.90 m high secondary collector storage tank from where it will be distributed to the respective flotation cells by dedicated metering pumps.

The proposed frother based on the metallurgical test program is MIBC and Anionic 919 flocculant will be required for concentrate thickening.

Reagent consumption estimates are included in Section 18.11.

18.6 Recoverability

Recoverability estimation is covered in Section 16.

18.7 Markets

This technical report has relied upon Taseko Mines Limited market and logistics experience at Gibraltar, a 2007 report by Neil S. Seldon & Associates Ltd. (NSA) on marketing, logistics, and penalty schedules, and a variety of publicly available market related information in its preparation.

Taseko has had discussions with a number of smelter groups regarding Prosperity concentrate.

Copper

Today there is every indication that copper prices have moved up to a higher long-term plateau. The key period for Prosperity is 2010-2020 which will cover the payback period.

The copper price as of October 2009 was at US$2.98/lb with the 3-year trailing average 2006-2009 copper price being US$2.91/lb.

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The forward 27-month and 63-month copper price curve as of December 10, 2009 was US$3.12/lb and US$3.04/lb, respectively

Current long-term mean consensus pricing for copper is approximately 1.85/lb.

As a reserve basis, a copper price of US$1.65/lb has been selected.

Gold

In US dollar terms along with other commodities, a higher price plateau seems assured. Production costs are rising, gold has found favour and many analysts are forecasting long term prices at or close to 2009 levels.

The gold price as of October 2009 was at US$1,040/oz with the 3-year trailing average 2006-2009 gold price being US$825/oz.

The forward 24-month gold price curve as of December 10, 2009 was US$1,152/oz.

Current long term mean consensus pricing for gold is approximately US$815/oz

As a reserve basis, a gold price of US$650/oz has been selected.

Silver

As with gold, in US dollar terms along with other commodities, a higher price plateau seems assured. Silver has found favour with investors and speculators and many analysts are forecasting long-term prices at or close to 2009 pricing.

The silver price as of October 2009 was at US$16.25/oz with the 3-year trailing average 2006-2009 silver price being US$14.23/oz.

As of December 10, 2009 the forward 24-month silver price curve was US$17.39/oz.

As a reserve basis, a silver price of US$10.00/oz.has been selected.

Smelter Terms

Based on their own review, involvement, and knowledge of the concentrate market and smelter terms, Taseko have assumed smelter charges at:

	TC US$80.00 per dmt
	RC US$0.08 per pound
	No Price participation at US$1.65/lb

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A summary of long-term assumptions for treatment & refining and other commercial terms for copper concentrates are presented below:

Payable Metals

Copper:		Deduct 1 unit and pay for balance of content with refining charges of US$0.08/lb
Silver:		Pay 90%, with a refining charge of US$0.45/oz
Gold:		Pay 97.50%, with a refining charge of US$6.00/oz

Deductions

Treatment Charge (TC):			CIF FO main Asian port parity, US$80/dmt
Price Participation (PP):			None at US$1.65/lb

Penalties:

Arsenic	US$3.00 per 0.1% over 0.1%
Antimony	US$3.00 per 0.1% over 0.1%
Mercury	US$0.20 per ppm over 20 ppm

Marketability

Prosperity concentrate quality is relatively low copper at 23% - 25%, relatively high gold at 35-45 grams/tonne and modest silver at 80 to 100 grams/tonne. This low copper means there will probably be some limit on the quantity that any one smelter will take as the grade is below the average smelter blend and reduces the metal output from the furnaces. However, all other considerations apart, a concentrate with this level of copper, gold and silver should be readily acceptable.

A consideration with Prosperity concentrate is the levels of arsenic, antimony and mercury when taken together. Individually, the arsenic in the range of plus/minus 0.2%, while above most smelter blends of 0.1%, will be acceptable.

Antimony in this concentrate at 0.3% -0.4% is well above the penalty threshold of 0.1% and while this is unlikely to affect the ability to sell, it will incur penalties and could well reduce individual smelters quantity interest.

Of these three penalty elements, mercury at 80 to 150 ppm, is probably the most significant. Mercury will incur penalties and not all smelters, even if they blend, will be prepared to take such quality.

From a logistic point of view the likely market for a major part of the production should be Asia. However, discussions with smelters and consideration of financing options may result in other preferable markets.

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In summary Prosperity concentrate will find a market but with the prospective quantity of about 200,000 tons per annum and given the penalty considerations will likely need to need be spread around a number of smelters.

Logistics

For overseas markets concentrates will be moved by truck to an expanded facility at Gibraltar’s existing load-out facility near Macalister, just north of McLeese Lake, B.C and then railed to the port facilities in North Vancouver or alternatively railed directly to Eastern Canada.

Transportation cost estimates for this study, based on transport to Asian markets are summarized below on a wet metric tonne (wmt) basis:

Truck mine to rail:	CDN$	28.00
Rail Freight to Vancouver:	CDN$	28.00
Concentrate Storage and Ship Loading:	CDN$	22.50
Ocean Freight and Sampling:	US$	60.00

Other Offsite Costs

Handling losses are estimated at 0.5%.

Marine insurance cover from the port of loading to the discharge port is assumed at a rate of 0.02% of concentrate NIV.

18.8 Contracts

No mining, concentrating, smelting, refining, transportation, handling, sales and hedging and forward sales contracts or arrangements have been negotiated to date. Rates and assumptions used within this study’s economic analysis follow industry norms.

18.9 Environmental Considerations

Environmental Assessment and Permitting

Background

The proposed Prosperity Project entered the provincial EA process in 1995. During the 1990s the EAO convened technical meetings of a Project Committee to discuss the information needs of government agencies and First Nations and developed the Project Report Specifications (Terms of Reference). In 2000 Taseko suspended the review of the Project to await more favourable economic conditions.

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In 2006 the review re-commenced and in February 2007 the federal Department of Fisheries and Oceans (DFO) requested that the federal review be referred to a Federal Review Panel. Despite over a year of consultation and discussion of joint federal/provincial panel agreement models by the Environmental Assessment Office (EAO) and the Canadian Environmental Assessment Agency (CEAA) it was not possible to reach agreement and in June 2008 the provincial Minister of Environment ordered that the provincial EA be undertaken by the EAO.

Federally, an EA of a proposed project is required under the Canadian Environmental Assessment Act, as amended, if a federal authority would be required to exercise certain powers or perform certain duties or functions in respect of a project for the purposes of enabling the proposed project to be carried out, in whole or in part. The Prosperity Project requires authorizations by Department of Fisheries and Oceans under the Fisheries Act, by Natural Resources Canada under the Explosives Act, and by Transport Canada under the Navigable Waters Protection Act.

On January 19, 2009, the federal Minister of the Environment appointed a Federal Panel and issued the three Panel members with Terms of Reference.

Prosperity Project Environmental Impact Assessment Report

On January 26th 2009 Taseko submitted an Environmental Impact Assessment Report to both the provincial EAO and the federal Panel.

Based on the October 2007 reserve, the project described in the Environmental Impact Assessment Report is for a project that will mine 487M tonnes of ore from the more than 1B tonnes of measured and indicated resources in the deposit. The proposed Project would have a mill throughput of 70,000 tpd and an active pit life of 17 years.

The Prosperity Environmental Assessment Report represents a compilation of expert opinion supported by scientific data and technical analysis representing more than a decade of information gathering, examination and engineering and scientific effort. Aspects of the physical environment including the atmospheric, acoustic, surface water hydrology and hydrogeology were examined and assessed in detail. It was concluded that in all of these areas there were no significant residual effects on the environment.

Following a thorough examination of the water and aquatic ecology, terrain and soils, vegetation and wildlife biotic environment it was concluded that the overall mine design, in combination with the implementation of regulatory standards, industry standard management practices and project-specific mitigations where needed achieves a development that will have no significant residual effect on the environment.

All mining activity is focused in the Fish Creek watershed. Placing all mining related infrastructure in this one watershed enables effective environmental management and reduces the risk to water quality and the salmon fisheries of the adjoining Taseko River. The unavoidable impact of the project on the trout fishery in Fish Lake will be offset by an extensive fish and fish habitat compensation plan developed in consultation with the provincial Ministry of

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Environment. The compensation plan involves the creation of a new lake, Prosperity Lake, and will ensure that the genetic integrity of the Fish Lake stock is preserved, that the recreational and First Nations fisheries opportunities that currently exist are maintained and that there is no loss of the productive capacity of the habitat that supports the fisheries resource.

Consideration was given to potential economic, social, community and human health issues including the effect of the project on other resources and their uses. It was concluded that while any negative effects on human health and other resource use are predicted to be not significant, the economic and socio-economic effects are considered to be positive. Project spending will stimulate both employment and business development which in turn will generate incremental income streams for government.

Taseko undertook an extensive and detailed archaeological impact assessment at the proposed mine site area and reported the findings from this study in the Impact Assessment Report. The information gathered indicates that the area was used for a range of activities including hunting, fishing, plant gathering and processing. Before the mine site is developed, mitigation measures designed to obtain a limited amount of additional information of cultural or heritage importance will be implemented.

Taseko began a First Nations engagement and consultation strategy in 1993 and it continues to this day. There are seven communities of Tsilhqot’in (Chilcotin) people and five communities of Secwepemc (Shuswap) people requiring consultation and engagement on the Project. Taseko’s engagement and consultation strategy has and continues to be developed and implemented in full collaboration with government. The Impact Assessment Report outlines and addresses all the key issues raised by First Nations.

Provincial Environmental Assessment

Early in the provincial EA process, the EAO and the CEA Agency agreed to coordinate the EA process to the extent possible to provide a single window for public participation and to minimize the potential for duplicate activity. The provincial and federal processes were coordinated for the review of the Terms of Reference and submission

During the June 2008 to March 2009 Pre-Application Review Stage the EAO established a working group (Working Group) comprised of local, provincial and federal government agencies and First Nations, held public meetings and comment periods and finalized Terms of Reference for the Environmental Impact Assessment Report. Final approved Terms of Reference were issued to Taseko by the EAO on January 9th 2009 and issued to Taseko by CEAA on January 19th 2009.

The review of The Application was commenced in March 2009 once the EAO accepted it as complete.

The EAO Assessment Report is to be finalized and forwarded on or before December 18th 2009 to Ministers for a decision. Ministers have until February 1st 2010 to decide on the issuance of an

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EA Certificate to Taseko Mines. Upon receipt of a provincial Project Approval Certificate, Taseko will proceed to permitting.

Federal Panel Review

On June 25th 2009 the federal Panel issued a Deficiency Statement to Taseko in which they outlined a number of areas where in their opinion, the information already provided was not sufficient to proceed to hearings. On August 18th 2009 Taseko submitted responses to the deficiency statement and following a 30 day public comment period the Panel determined that in all but two areas, aboriginal fisheries and First Nations cultural heritage, there was sufficient information to proceed to hearings. In their letter of October 6th 2009 the Panel gave First Nations until November 17th 2009 to provide Taseko with information on their current use of lands and resources for traditional purposes. Following receipt and review of Taseko’s assessment of this additional information it is anticipated that the Panel will announce a hearing schedule.

The Panel Terms of Reference indicate that best efforts should be applied to complete the hearings within a 30 day period. Following the close of hearings the Panel has 60 days in which to finish and submit its report to the federal Minister of Environment. Following a decision by federal cabinet, federal regulators can proceed to exercise their regulatory duties in accordance with a schedule as outlined within the Major Projects Management Office (MPMO) Project Schedule.

Current Status

A provincial government decision whether or not to issue the Environmental Assessment Certificate will be made by no later than February 1st 2009.

If approved, the provincial EA Certificate will be for the 487M tonne reserve, 70,000 tonnes per day mine plan outlined in the Environmental Impacts Assessment Report and not the 830M tonne reserve announced on November 2nd 2009 that is the subject of this report. Should at some later date Taseko Mines Limited decide to change the present mine plan, it would do so in a manner entirely consistent with both federal and provincial statutory requirements in effect at that time.

The federal Panel review process is ongoing and not expected to be concluded until at least the second quarter of 2010.

Closure and Reclamation

A conceptual reclamation plan has been provided as part of the environmental assessment. A more detailed reclamation plan will be required as part of the BC Mines Act mine permit application when it is submitted.

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The current conceptual reclamation plan includes reclamation planning information (maps and descriptions) corresponding to the stages of mine construction, 5-year mark, and mine closure, including interim reclamation objectives, proposed end land uses and the means by which reclamation work will achieve objectives. Upon mine closure, surface facilities will be removed in stages and full reclamation of the Tailings Storage Facility (TSF) will be initiated. General aspects of the closure plan include:

· Selective discharge of tailings around the facility during the final years of operations to establish a final tailings beach that will facilitate surface water management and reclamation;

· Dismantling and removal of the tailings and reclaim delivery systems and all pipelines, structures and equipment not required beyond mine closure;

· Construction of an outlet channel/spillway at the east abutment of the Main Embankment to enable discharge of surface water from the TSF to the open pit and ultimately to Lower Fish Creek. This full closure scenario will also work well in the unlikely event of premature permanent closure of the mine;

· Removal of the seepage collection system at such time that suitable water quality for direct release is achieved;

· Removal and regrading of all access roads, ponds, ditches and borrow areas not required beyond mine closure;

· Recontouring surfaces to facilitate optimum plant production, appropriate site drainage, and animal access;

· Site preparation to alleviate compaction and facilitate drainage;

· Soil replacement to stimulate plant establishment, suitable quality forage for animal consumption, and long-term sustainable ecosystem function;

· Seeding and planting areas as soon as possible after placement of soil with a seed mix suitable for erosion protection and is consistent with end land use goals;

· Long-term stabilization of all exposed erodible materials.

Permitting

Federal and provincial licenses and permits that may be required by the Project, including requirements associated with any necessary amendment to Schedule 2 of the Metal Mine Effluent Regulations may include most (if not all) of the following:

· BC Ministry of Energy, Mines and Petroleum Resources - Permit Approving Work System and Reclamation

· BC Ministry of Agriculture and Lands - License of Occupation for Water Diversions/ Discharge Lines, Borrow/Gravel Pits, Staging Areas During Construction; Statutory Right of Way for Transmission Line

· BC Ministry of Forests and Range - License to Cut

· BC Ministry of Environment - Water Licenses for water Storage, Diversion, and Use; Approvals for changes in and about a stream; Waste Management Permit for Effluent discharge (sediment, tailings & sewage), Air discharge (crushers,

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concentrator), Refuse; Fish Culture Permit for Transport and possession of live rainbow trout

· BC Ministry of Health - Construction Permit for construction, alteration or extension of a water works system; Operation Permit for water works system; Camp Operation Permits - Drinking water, sewage disposal, sanitation and food

· BC Ministry of Tourism - Alteration Permit for disruption of archaeological resources

· Fisheries and Oceans Canada - Section 35(2) Authorization Fish Habitat Compensation Agreement

· Environment Canada - Schedule 2 Amendment

· Canadian Coast Guard - Navigable Water: Stream Crossings Authorization

· Natural Resources Canada - Explosives Factory License; Explosives Magazine License

· Transport Canada - Approval for Ammonium Nitrate Storage Facilities

· Industry Canada - Radio Licenses

· Canadian Nuclear Safety Commission (Natural Resources Canada) - Radioisotope License

18.10 Taxes

The economic model was run on a before tax basis. BC mining taxes were estimated and included in the cash flow model. The project will also be subject to Federal and Provincial income taxes but these rates are not fixed and it is believed that tax planning methods will be available to minimize the affect on project economics.

18.11 Capital and Operating Cost Estimates

Capital Cost — Summary

The direct and indirect capital cost to bring the Prosperity Project into production at a design ore throughput rate of 70,000 tonnes/day is estimated to be $814 million as summarized in Table 18-7.

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Table 18-7

Pre-production Capital Cost ($ x 1000)

Description	CDN$(x1000)

Site Preparation	$	9,654
Mining	$	68,356
Crushing, Conveying & Stockpiling	$	51,741
Concentrator	$	227,125
Tailings Disposal & Reclaim Water	$	21,914
Site Infrastructure	$	95,539
Offsite Infrastructure	$	55,199
Total Direct Costs	$	529,529
Total Indirect Costs	$	161,532
Owners Costs	$	19,585
Contingency	$	103,659

TOTAL PROJECT COSTS	$	814,304

The capital cost estimate is expressed in constant 2nd quarter 2009 Canadian dollars using exchange rates of:

$US:$CDN		0.95
$EURO:$CDN		0.65

Indirect capital costs forming part of the overall capital cost estimate for the project have been estimated to be $162 million. A summary of the indirect cost estimate by area is shown in Table 18-8.

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Table 18-8

Pre-production Indirect Costs ($x1000)

Description		CDN$(x1000)

EPCM		51,138
Construction Indirects		64,745
Catering & Camp Maintenance		15,843
Vendor Reps		3,744
Critical Spares		11,859
First Fills		3,773
Freight and Transport		10,430
TOTAL INDIRECTS		161,532

The sustaining and additional equipment capital cost estimate is $701 million over the operating life-of-mine and is summarized in Table 18-9.

Table 18-9

Life of Mine Sustaining Capital Cost ($ x 1000)

Description	CDN$(x1000)
Additional Mining Equipment	$	248,037
Equipment Replacement	$	273,642
Tailings	$	112,830
Concentrator and Infrastructure	$	29,752
Closure	$	36,484
TOTAL SUSTAINING CAPITAL	$	700,745

Capital Cost Schedule

The implementation schedule from issuance of all required permits to commissioning is approximately two years. The expenditure of capital has been scheduled as 35% in year -2 and 65% in year -1.

Equipment addition and sustaining capital expenditures for the Prosperity Project were estimated on the basis of equipment requirements, equipment life, dam construction requirements, and a closure accrual based on copper production over the operating life of the mine.

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Basis of Estimate - Direct Costs

The capital cost estimates include all direct costs, indirect costs and contingency to construct all facilities required to bring the Prosperity mine and concentrator to full production. Major pit production equipment has been assumed to be leased with appropriate down payments and annual payments prior to concentrator commissioning captured as capital and subsequent payments included in sustaining capital. Owners cost allowances are included “below the line” (i.e. do not attract a contingency allowance and are not included under “indirects”)

The following technical documentation forms the basis of the engineered plant solution:

· Process Design Criteria

· Process Flowsheets;

· Sizing of all major equipment items.

· Mechanical Equipment List;

· General Arrangements/Layouts/Plot Plan;

· Electrical Equipment List;

· Electrical Load List;

· Civil Drawings;

These documents provide the basis for pricing of the complete works. Accordingly the direct cost estimate was derived from the following main input categories:

· Equipment Quotations

· Bulk Material and Earthworks Takeoffs

· Estimates by third parties

· Quotations on standard structures

· Labour & other unit rates developed

Input to the direct cost estimate was collected during the height of mining equipment and general construction demand in mid-2008. In order to reflect 2009 construction and mining market conditions, Taseko has applied reasonable factors based on current purchasing and construction experience at Gibraltar and decreases in base commodity prices and labour rates relative to mid 2008 in estimating the direct capital cost in 2009.

Mine pre-production and life of mine plan units, costs, and equipment requirements were developed by AKF Mining Services Inc, and Taseko Mines Limited.

Based on the above studies, mining fleet requirements were developed and a mining equipment list finalized. Budget quotations and current Gibraltar pricing for this equipment forms the basis of the mining fleet cost estimate.

The engineering and estimating of the tailings, coffer dams, water diversion and other water management systems as well as the waste rock designs were done by Knight Piesold.

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Knight Piesold provided the tailings embankment construction material requirements, which were incorporated into the pit plan to utilize waste material from the pit to construct the embankments. They also provided the cost estimates for embankment material placement while Taseko provided the pre-production tailings embankment construction material delivery costs.

The estimate also allows for tailings discharge and water reclaim infrastructure including barge, pumps, tailings pipeline and water return water lines.

Engineering and cost estimating for the overhead transmission line from Dog Creek to Prosperity was done by Ian Hayward International Ltd. The estimate for the switching station at Dog Creek was completed by SNC Lavalin for BCTC/BC Hydro.

Labour rates and employment conditions for remote site construction in Northern British Columbia were provided by the Christian Labour Alliance of Canada (CLAC). This organization provides qualified crafts for open site projects throughout Canada

The construction work week used for the estimate is six ten-hour days per week. Overtime over eight hours per day or forty hours per week is paid at time and one-half. The turnaround cycle is three weeks in, and one week out. The estimate includes travel compensation per turnaround. Travel pay is not included.

The estimated direct labour rates used are crew composites for each commodity and include general foremen, foremen, lead hands, journeymen and apprentices. The cost elements include:

· Base Wage Rate

· Premium overtime

· Fringe Benefits;

· Government Assessments;

· Payroll Service charge at 3% of payroll;

· Small tools and consumable supplies at $2.50 per hour worked;

· Contractor Home Office Overheads at 10% of total; and

· Contractor profit at 7.5% of total.

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Basis of Estimate - Indirect Costs

The indirect cost estimate for the project includes the following items:

· EPCM based on the proposed project staffing plan, including uplift and completion bonuses, office expenses, computer fees, travel costs, and the use of additional consultants;

· Construction indirects based on built up allowances for all required services;

· Catering and camp maintenance;

· Diesel generators for construction and camp power during the construction period;

· Initial fills;

· Vendor representatives;

· Spare parts;

· Freight

Input to the indirect cost estimate was collected during the height of general construction demand in mid-2008. In order to reflect 2009 construction market conditions, Taseko has applied reasonable factors based on current construction experience at Gibraltar and decreases in base commodity prices, increases in equipment availability, and labour cost reductions relative to mid 2008 in estimating the indirect costs in 2009.

Basis of Estimate - Owners Costs

The owner’s costs were developed by Taseko to cover all owners costs excluding the mine fleet and mine preparation during the construction period. They are derived from time-based resource allocations and include:

· Salaries for owners team, pre-production supervision and support staff, and ramp up to full staffing for the beginning of year 1

· Recruitment and relocation

· Temporary accommodation, transportation, and travel

· Williams Lake office and supplies

· Insurance

· Community outreach

· Environment and assaying

· Legal costs

· Marketing

· Permits, licenses, and leases

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Contingency Estimate

Contingency is the sum of money added to an estimate to cover the cost of unforeseeable occurrences after the letter and intent of the scope described in the feasibility report have been duly and diligently identified, quantified and costed or otherwise provided for. Examples of contingent items include but are not limited to:

· Estimate errors and omissions;

· Design developments;

· Pricing variations;

· Unusual weather excluding extreme consequences of extreme events

The contingency capital is estimated to be 15% of the direct plus indirect capital cost, amounting to $103.7 million.

The contingency allowance is intended as a measure of the level of accuracy which can be placed on the capital cost estimate to account for unforeseen costs within the scope of the estimate as well as for unforeseen construction schedule accelerations and delays. Contingency costs may also be incurred due to undefined items of work or equipment beyond the control of the builder, or to uncertainty in some quantity estimates or unit prices for labour, equipment and materials. The contingency allowance should be expected to be spent in the normal course of events.

Sustaining Capital

The sustaining capital cost estimate is expressed in constant 2nd Quarter 2009 Canadian dollars.

The cost of additional and replacement mining and mining support equipment has been based on the replacement and purchasing schedules developed by Taseko utilizing budgetary quotes and Gibraltar actual equipment purchase prices.

The ongoing cost of plantsite mobile (ancillary) equipment replacement has been based on the replacement and purchasing schedule developed by Taseko.

An additional annual allowance for sustaining capital related to non-mobile equipment replacement has been calculated to achieve a minimum annual value of $0.07/tonne mined for years 1 through 24, decreasing annually to $0.01/tonne mined in year 30, with no additional allowance in the final three years. This allowance has been allocated to the mill.

The first 16 years of tailings embankment construction, instrumentation, pipework, reclaim water systems and seepage control has been estimated by Knight Piesold. For the year 16 to year 33 period Taseko has extrapolated a constant cost per tonne milled based on the detailed build-up in first 16 years.

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Capital Cost Exclusions

The following costs were not included in the capital cost estimates:

· Environmental, archaeological and ecological considerations, other than those incorporated in the current design;

· Costs for acquisition of Rights-of-Way;

· The cost of producing any environmental impact statement and obtaining environmental permits and approvals from local or national authorities;

· Financing charges and interest;

· Currency exchange fluctuations after Sept 1, 2009;

· All costs associated with weather interruption of construction operations;

· Costs of Public Relations activities and any costs of impacts to construction work associated with implementation of Public Relations operations;

· Escalation beyond second quarter 2009;

· Price fluctuations due to unusual market conditions;

· Provision to attract and retain qualified labour during construction;

· Value added tax;

· Owner’s head office costs;

· Exploration expenses;

· Construction reclamation costs;

· Sunk costs;

· Federal goods and services taxes; and

· Import duties.

Operating Cost Estimate - Summary

The base case operating costs for the Prosperity Project, including mining, milling, and general and administrative costs have been estimated in 2nd quarter 2009 Canadian dollars and include no allowance for escalation or exchange rate fluctuations.

The average project site operating cost for the 33 year life-of-mine is estimated to be $7.51/t of ore milled.

The operating costs are presented in three major segments:

· Mining: includes the direct costs of mining, including drill, blast, load, and haul activities, roads and dumps, and a general mine expense

· Milling: includes all operating costs associated with the concentrator from the dump pocket of the gyratory crusher to the discharge of tailings into the tailing pump box and to the loading of concentrate trucks at the mill site. It also includes the power costs for all non-pit activities, plant services labour and equipment, and general site infrastructure and buildings.

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· General and Administrative: includes salaries and wages of administrative personnel including purchasing and warehouse. It also includes all fixed costs relative to administration, warehousing, employee relations, IT services, safety and security, training, and the operations camp.

The estimated life of mine average unit costs for each major operating cost area are shown in Table 18-10.

Table 18-10

Life-of-Mine Unit Costs

Operating Category	Operating Cost ($/tonne ore processed)
Mining	$	3.14
Milling	$	3.85
General and Administrative	$	0.52
Total	$	7.51

Operating Cost Basis of Estimate

Taseko developed specific operating cost estimates for mining, milling, and general administration. Other key consultants and services providers contributing operating cost estimates include:

·	Gibraltar Mines Ltd. (Gibraltar)	Mining
·	Knight Piesold Consulting ( Knight Piesold)	Pit Dewatering
		Pit Wall Depressurization
		Tailings Dam Material Placement

The operating cost estimates are based on the following general project data:

· Pit design, phasing and scheduling using appropriate haulage profiles and equipment cycle times;

· Budgetary quotations, current Gibraltar pricing and long-term Gibraltar pricing assumptions for major consumables including power, grinding media, reagents, mill and crusher liners, fuel, tires, and explosives;

· Process reagent consumption rates generated from metallurgical test work;

· Power requirements from material test work; and

· Power consumption requirements generated from an assessment of the mechanical equipment and service electrical loads.

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· Pit water inflows and depressurization requirements from geotechnical and hydrogeological studies

Power unit cost was based on Gibraltar Mine’s current cost. Mill and infrastructure power consumption was derived based on the connected load data from the mechanical equipment list. Energy consumption was based on a 94% motor efficiency. The estimated average annual energy consumption for operating years 7 to 17 (after pumping of tailings has reached full power consumption) is 700 GWh. An allowance for pit related power totals an additional 50 GWh for purposes of demand.

Operating labour wages have been based on the most recent labour contract negotiated with the Construction and Allied Workers’Union (CLAC), Local 68 at Gibraltar Mines Ltd. reflecting 2009 labour rates.

Wages include payroll burdens to cover employer costs for employee benefits, holiday pay, Canada Pension Plan contributions, Worker’s Compensation assessments, employment insurance premiums, and life and long term disability insurance premiums.

Salaries for management, supervisory staff, and technical staff have been based on 2009 Gibraltar salary levels. Salaries include payroll burdens to cover employer costs for employee benefits, holiday pay, Canada Pension Plan contributions, Worker’s Compensation assessments, employment insurance premiums, and life and long term disability insurance premiums.

A breakdown of operating manpower by year is summarized in Table 18-11.

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Table 18-11

Summary of Estimated Annual Average Operating Manpower (Years 10-25)

		Years 10-25
Mining
Supervisory and Technical Staff		49
Mining Operations		159
Mining Maintenance		76

Subtotal		284

Milling
Staff		20
Mill Operations		45
Mill Maintenance		54
Plant Services		25

Subtotal		144

General & Administration
Administration Staff		22
Purchasing and Warehouse		9
Warehouse Hourly		4

Subtotal		35
Operating Manpower Total		463

Supplies and consumables costs for mining operations include such items as explosives, blasting accessories, fuel, oil and lubricants, filters, tires, equipment wear parts, equipment mechanical and electrical component replacement parts. Costs for explosives, fuel, drill steels and bits have been based on current Gibraltar pricing and contracts, and long range forecasts. Mechanical and electrical component parts including overhaul costs are based on supplier estimates of frequency and cost.

The replacement of parts due to normal wear and tear and equipment breakdown was considered part of the operating cost. Periodic replacement of capital equipment was considered sustaining capital and not included in the operating costs. Initial stocking of spare maintenance parts and operating supplies was considered to be initial capital and therefore has not been included in the operating cost estimate.

Supplies and consumables cost for mill operations include all major items consumed during the operating of the concentrator. Consumption rates for consumable supplies are based vendor recommendations, metallurgical test work, and Gibraltar operating experience. Unit prices for consumable items are based on Gibraltar current pricing or budget quotations from suppliers.

Annual non-pit mechanical maintenance supplies, electrical maintenance supplies, and building maintenance costs are based on percentages of initial capital costs.

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Mining Costs - Summary

The life-of-mine unit mining operating cost is estimated at $3.14/t ore milled or $1.27/t material mined. This operating cost is referenced to the 70,000 tpd ore production schedule over a plant operating life of 33 years

The estimated life-of-mine average mining unit cost is summarized in Table 18-16.

Table 18-12

Life-of-Mine Direct Mining Unit Cost

Area	$/t milled	$/t mined
Drill	0.22	0.09
Blast	0.37	0.15
Load	0.25	0.10
Haul	1.48	0.60
Roads and Dumps	0.34	0.14
General	0.48	0.19
Total	3.14	1.27

Processing Cost Summary

The average processing cost over the life of the mine including processing of ore stockpiles is estimated to be $3.85/t of ore.

The battery limits for operating costs associated the concentrator are from the dump pocket of the gyratory crusher to the discharge of tailings into the tailing pump box and to the loading of concentrate trucks at the mill site. The concentrator area also captures the power costs for all non-pit activities as well as the plant services labour and equipment.

Concentrator operating costs have been separated into 6 categories; labour, power, consumables, maintenance, mill general and plant services. They have been calculated on an annual and a cost “per tonne of ore processed” basis. The operating costs vary according to grinding media consumption, the requirement to pump tailings, and varying manpower levels.

Typical annual concentrator operating costs by category are shown in Table 18-13.

Processing costs include plant services which encompass tailings dam construction and general site infrastructure support. The plant services component of the milling cost includes only labour and equipment operating costs as items such as power and infrastructure materials are captured within the mill maintenance costs.

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Table 18-13

Typical Annual Processing Operating Cost Summary (Year 7-25)

	Annual Cost	Unit Cost
Area	($x1000)	($/t milled)
Labour
Staff	2,204	0.086
Mill Operations	3,893	0.152
Mill Maintenance	5,338	0.209
Plant Services	2,291	0.090
Subtotal	13,726	0.537
Power
Water Systems including Water Storage & Distribution	309	0.012
Fuel Systems	16	0.001
Crushing	891	0.035
Conveyors and Stockpile Reclaim	159	0.006
Concentrator Building including HVAC	926	0.036
Grinding	15,199	0.595
Pebble Crushing	133	0.005
Flotation and Air Systems	2,713	0.106
Regrind	2,000	0.078
Concentrate Thickening and Loadout	144	0.006
Concentrator Reagents	43	0.002
Procces Water	216	0.008
Tailings Pumping	503	0.020
Water Supply	1,122	0.044
Eng $ Admin. Facitlites	132	0.005
Mine Service Facilities	55	0.002
Assay Laboratory	97	0.004
Operations Camp	372	0.015
Plant distribution Losses	1,365	0.053
Subtotal	26,396	1.033
Consumables
Grinding Media - SAG Mill Balls	8,162	0.319
Grinding Media - Ball Mill Balls	20,399	0.798
Grinding Media - Regrind Mill Balls	6,874	0.269
Primary Collector- Xanthate	2,557	0.100
Secondary Collector	2,459	0.096
Frother	1,345	0.053
Flocculant	35	0.001
Lime	3,252	0.127
Subtotal	45,083	1.763
Maintenance
Concentrator Reagents	6,025	0.236
Concentrator Reagents	4,916	0.192
Concentrator Reagents	1,162	0.045
Subtotal	12,104	0.473
Mill General
Buildings	426	0.017
Subtotal	426	0.017
Plant Services
Mobile equipment	730	0.029
Subtotal	730	0.029
TOTAL COST	98,466	3.852

The power cost allocated to the mill includes all non-pit related power. The estimated average annual energy consumption for operating years 7 to 18 (after tailings pumping is established at normal annual load) is 700 GWh.

144

The average power consumption per tonne of ore processed of the SAG mills and ball mills will increase as mining progresses to depths where the hardness of the ore is geologically related to the gypsum line. Since the mills are expected to operate at full power, the annual power consumption in the grinding circuit is expected to remain unchanged independent of the ore hardness.

Reagent consumption rates are based on the results from metallurgical test work. There may be some justification for applying slightly lower consumption rates in a full scale operation than those experienced in the laboratory due to the use of reclaim water in the full scale operation which will contain some usable reagents and also due to reagent optimization that will occur in the full scale operation. However the reagent consumption rates experienced in the metallurgical test work have not been reduced for use in this operating cost estimate.

General and Administration Costs

The estimated average life of mine cost for general and administrative (G&A) mine functions is estimated to be $0.52/t ore milled. Major categories are summarized in Table 18-14.

Administrative salaries and wages are based on an estimated G&A manpower complement while salary and wage rates including burdens have been estimated from rates at Gibraltar.

The major fixed costs in this area include property assets insurance, taxes, freight, bussing and environmental costs.

Property assets insurance costs have been based on a loss limit of $150,000,000 for any one loss with a $250,000 deductible.

Return personnel transportation from Williams Lake to the mine will be provided by a chartered bus service. The transportation cost for all operating personnel has been estimated at $500,000 annually.

An annual allowance of $350,000 has been made for environmental services.

Other components of fixed G&A costs have been based on current Gibraltar costs.

145

Table 18-14

Estimated Typical Annual General & Administration Costs Years 10-25, ($000’s)

Year		10 - 25

Salaries and Wages
Administration		1,897
Purchasing and Warehouse		1,151
Subtotal		3,048

Fixed Costs
Administration		1,829
Warehouse		545
Employee Relations		762
Computer Services		323
Safety and Security		436
Training		216
Environment		350
Subtotal		4,460

Camp Costs		6,049
Subtotal		6,049

Total		13,557

Housing for both site operating and ongoing contract personnel will be provided at the Prosperity camp. Camp costs at a rate of $70/man-day have been estimated and are accounted for in the G&A area. Catering costs include the costs associated with the housing of ongoing contracted personnel for outside services such as mining explosives supply, pit dewatering, horizontal drilling, mining mobile equipment erection, and guests. Costs are inclusive of camp management, maintenance, housekeeping and catering.

Working Capital

An amount equivalent to 20% of first full year project operating costs (approximately $52.0 million) was charged against the project cash flow in Year 1 to provide for the production of initial product inventories. This was credited back to the project cash flow in Year 33 to reflect product inventory drawn down and recapture of receivable accounts.

146

18.12 Economic Analysis

Summary

The metal prices used in this economic analysis are US$650/oz gold, US$10.00/oz silver, and US$1.65/lb copper as outlined in Section 18-6. These are conservative values well below current long term forecast prices and are used solely as the basis for determining the mineral reserve. Applying long term forecast prices will result in significantly greater economic performance metrics.

Unless otherwise stated all dollar amounts used in the analysis are in constant 2nd quarter 2009 CDN$.

An exchange rate of $0.82 USD per CDN$ has been used. Inflation factoring has not been applied.

The analysis is based on an ore reserve that processes 831,218,000 tonnes grading on average 0.23% copper and 0.41 g/t gold over a 33 year mine life, and a mill production rate of 25,560,000 t/y (70,000 t/d).

Key indicators derived from an economic analysis of the project at reserve metal prices only are summarized in Table 18-15.

Table 18-15

Key Reserve Indicators

Total copper production		1,655,000 tonnes/3,648,000,000 lbs
Total gold production		7,720,000 troy oz
Total silver production		19,800,000 troy oz
Mine Life		33 years
Pre-tax return on investment (ROI)		10.0%
Pre-Tax payback period		8.0 years
Pre-production Capital Cost		$814,000,000
LOM Sustaining Capital		$701,000,000
LOM Average site operating costs		$7.51/t ore processed
LOM Average off property costs		$3.06/t ore processed
LOM average cost (net of by-product credits)*		US$0.59 per lb Cu produced

*Total site and off property costs including all TC/RCs and transportation

147

The reserve has been evaluated on a “stand-alone” basis assuming one owner with 100% equity financing for the project and no external corporate structures.

Exchange Rate

The average monthly US/Can dollar exchange rate has varied from approximately 0.80 to par over the 2005 — 2009 period. It is accepted that strong commodity prices lead to strength in the Canadian dollar against its US counterpart. The value of the Canadian dollar is susceptible to commodity prices, particularly oil and metals and it is reasonable to link prediction of the US/Can dollar exchange rate to the commodity cycle.

The correlation between the value of the Canadian dollar in terms of the US/CDN dollar exchange rate and the price of copper is very evident as depicted in Figures 18-10 and 18-11.

Figure 18-10

Constant 2008$ Copper Correlation with CDN/US Currency Exchange

148

Figure 18-11

Historical Correlation — Copper and Exchange Rate 1989 to 2008

This reserve update, using long term copper prices of US$1.65/lb reflective of the commodity cycle has used an exchange rate of US$0.82:CDN$1.00.

Production Schedule

The Prosperity mine will produce an estimated 25,560,000 tonnes/year (70,000 t/d) of ore by open pit mining methods with a strip ratio of 1.5:1. An estimated 60Mt of ore will be stockpiled in the first five years and processed prior to the end of operations. The economic model is based on a copper recovery to a single copper-gold concentrate with life-of-mine copper and gold recoveries of 87% and 69% respectively using conventional flotation methods.

Over the mine life, a total of approximately 227,000 tonnes of concentrate (wet basis) will be produced annually containing a life-of-mine average of 24% copper, 35 g/t gold and 89 g/t silver. Concentrate will be trucked to Gibraltar’s concentrate handling facilities for transfer to rail transport to various points of sale; primarily through the Port of North Vancouver for shipment overseas.

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Revenue

The project’s NSR in this economic analysis has been calculated using price forecasts for copper, gold, and silver, concentrate smelter and penalty terms, and inland and ocean freight costs as outlined in Table 18-16.

Table 18-16

NSR Assumptions

Gold		US $650 per troy ounce
Copper		US $1.65 per lb
Silver		US $10.00 per troy ounce
Exchange Rate		US $0.82/CDN $
Treatment Charge		US $80/dmt
Copper Refining Charge		US $0.08/lb
Silver Refining Charge		US $0.45/oz
Gold Refining Charge		US $6.00/oz
Copper Payable		96.5%, minimum 1 unit
Silver Payable		90%
Gold Payable		97.5%
Penalties (Sb, As, Hg)		US $30.20/dmt
Moisture		7.5%
Mine to Port		CDN $61.00/wmt (years 1-5), CDN $56.0/wmt after year 5
Port Charges		CDN $22.50/wmt
Ocean Freight/Assaying		US $60/wmt
Losses		0.5%
Insurance		$0.02/$100 NIV

Deductions from revenue have been made for the presence of mercury, arsenic and antimony that exceed specified thresholds in the concentrate as follows:

· Arsenic - A penalty of US$3.00 for each 0.1% in excess of 0.1%

· Antimony - A penalty of US$3.00 for each 0.1% in excess of 0.1%

· Mercury - A penalty of US$0.20 for each 1ppm Hg in excess of 20 ppm

Estimated impurity content in concentrate varies with depth, with penalties decreasing from US$37.70/dmt in the upper reserve to US$16.50/dmt in the lower. US$30.20/dmt is the estimate for the middle zone and this value has been used in this analysis.

Operating Costs

Operating costs were estimated in detail and are presented in Section 18.10. The operating costs

150

for mining, milling, and general and administration are indicated in Table 18-17.

Table 18-17
Operating Unit Costs

Operating Category	Operating Cost ($/tonne ore processed)
Mining	$	3.14
Milling	$	3.85
General and Administrative	$	0.52
Total	$	7.51

Capital Costs

The project capital cost estimate used in this economic analysis has been estimated in detail and presented in Section 18.10. Costs have been estimated based on feasibility level engineered designs, on quantity take-offs for construction materials, on CLAC construction labour rates, on Gibraltar labour rates, and on budget level quotations for equipment and purchased packages such as pre-engineered or modular structures.

Table 18-18 provides a categorization of the estimated project capital cost in CDN$.

151

Table 18-18

Capital Cost

Description	CDN$(x1000)

Site Preparation	$	9,654
Mining	$	68,356
Crushing, Conveying & Stockpiling	$	51,741
Concentrator	$	227,125
Tailings Disposal & Reclaim Water	$	21,914
Site Infrastructure	$	95,539
Offsite Infrastructure	$	55,199
Total Direct Costs	$	529,529
Total Indirect Costs	$	161,532
Owners Costs	$	19,585
Contingency	$	103,659
TOTAL PROJECT COSTS	$	814,304

Ongoing Capital Expenditures

Ongoing project capital expenditures have been estimated in detail and presented in Section 18.10. Ongoing capital expenditures will be required for additional mine equipment, ongoing replacement of mobile plant and mining equipment, staged pit de-watering and well installation, staged TSF embankment construction, and final property closure and reclamation.

Ongoing capital over the life-of-mine is estimated to be $701 million. No contingency has been applied to ongoing capital.

In order to account for reclamation and closure, an accrual of $0.01/lb copper produced has been included and reflected as a capital cost.

Working Capital

152

Taxes

No allowance has been made for Federal and Provincial income tax.

The only taxes calculated in this analysis are with respect to B.C. Mineral Taxes.

No allowance has been made for GST or provincial sales tax.

Life-of-Mine Cash Flow

A life-of-mine cash flow using reserve base metal prices of US$1.65/lb copper and US$650/oz gold details are presented in Table 18-19. The case cash flow demonstrates positive annual earnings and has been estimated using the following assumptions:

· $1.65/lb copper and US$650/oz gold

· Values are exchanged between currencies at CDN$1.00 = US$0.82

· Unless stated otherwise, all values are expressed in CDN$.

· Unless stated otherwise, all values are estimated in 2nd quarter 2009 dollars.

· No escalation is applied in the evaluation for inflation.

· The reserve is evaluated on a stand-alone basis. No external corporate structures are considered.

· The project is assumed to be financed on a 100% equity basis and 0% debt.

· Cashflow shown is calculated as gross revenue less total operating costs, capital, and mineral taxes.

Using only reserve basis metal prices and exchange, on a pre-tax basis the project shows a ROI of 10.0% and an estimated pre-tax payback period of 8 years.

153

Table 18-19

Reserve Basis Cash Flow

		Year		-2		-1		1		2		3		4		5		6		7		8		9		10		11		12		13		14		15		16
Net Cashflow		(000) $C		-289,281		-525,023		-23,319		163,394		189,579		129,315		145,367		83,891		55,225		74,135		29,822		34,638		13,673		58,213		58,266		75,414		83,930		115,684

		Year		17		18		19		20		21		22		23		24		25		26		27		28		29		30		31		32		33
Net Cashflow		(000) $C		102,879		37,857		26,346		47,158		49,568		44,900		53,860		74,476		101,547		166,545		184,314		221,922		251,729		278,250		184,749		63,880		109,051

154

Sensitivity Analysis

The relative sensitivity of the reserve to variations in revenue, capital cost and operating cost has been assessed by means of a “sensitivity analysis” which factors the above variables independently from 80% to 120% of their base case value.

The return on investment (ROI) sensitivity plot and table in Figure 18-17 describe the relative impact of changes to the following major economic variables, namely:

Site operating cost

Offsite costs

Initial capital cost

Sustaining capital cost

Exchange rate

Copper price

Gold price

ROI remains positive over a wide range of the major variables, demonstrating the stability of the reserve. The reserve is most sensitive to the currency exchange rate followed by operating cost, and metal prices.

155

ROI Sensitivity

Variable	-20%		-15%		-10%		-5%		Base		5%		10%		15%		20%

Site Op Cost	14.0	%	13.1	%	12.0	%	11.0	%	10.0	%	8.9	%	7.9	%	6.8	%	5.7	%
Offsite Cost	11.6	%	11.2	%	10.8	%	10.4	%	10.0	%	9.6	%	9.1	%	8.7	%	8.3	%
Initial Capital Cost	13.1	%	12.2	%	11.3	%	10.6	%	10.0	%	9.4	%	8.8	%	8.4	%	7.9	%
Sustaining Capital Cost	10.3	%	10.2	%	10.2	%	10.1	%	10.0	%	9.9	%	9.8	%	9.7	%	9.7	%
FX	2.3	%	4.3	%	6.3	%	8.1	%	10.0	%	11.8	%	13.6	%	15.4	%	17.3	%
Cu Price	5.4	%	6.6	%	7.8	%	8.8	%	10.0	%	11.1	%	12.2	%	13.3	%	14.3	%
Au Price	5.9	%	6.9	%	8.0	%	9.0	%	10.0	%	11.0	%	12.0	%	13.0	%	13.9	%

156

19. Interpretations and Conclusions

Location

The Prosperity Project is located 125 km southwest of the City of Williams Lake in the Cariboo-Chilcotin region of British Columbia, Canada.

Property and Access

The property is 100% owned by Taseko and is not subject to any royalties or carried interests. Production from the property will be subject to the Gold Stream Agreement. The mineral claims are currently in good standing until the year 2018.

The claims are situated within an area that was the subject of an aboriginal Rights and Title case in which the Supreme Court of B.C. recently found that the Tsilhqot’in Nation’s rights do not include the subsurface rights to the area around Fish Lake.

Access and infrastructure is adequate for the development of a large scale open pit operation with existing road access to the property, confirmed technical viability of hydroelectric power within 120 km, adequate water available, and rail load-out services close to Williams Lake.

Exploration

Taseko carried out ongoing and systematic exploration programs on the Project from 1991 — 1999, increasing drilling to 150,090 m in 470 holes, outlining a large porphyry gold-copper deposit. The Company and its consultants also carried out progressive engineering, metallurgical and environmental studies.

Taseko re-initiated environmental and engineering work on the Prosperity Project in late 2005. In 2007 approximately 1,800 m was drilled in 6 holes to collect sufficient material representative of the various alteration zones anticipated to be encountered in the first six years of mining.

Geology and Resources

The geology of this porphyry-type gold-copper deposit is well understood. The deposit is oval in plan and is approximately 1,500 m long, 800 m wide and extends to a maximum drilled depth of 880 m. Pyrite and chalcopyrite are the principal sulphide minerals in the deposit. They are uniformly distributed as disseminations, fracture-fillings and sub-vertical veinlets throughout the host volcanic and intrusive units in the deposit. Native gold occurs as inclusions in, and along microfractures with, copper-bearing minerals and pyrite.

Sampling, sample preparation, analysis and security meet industry standards. The results of the Taseko verification program indicate that the database is of good quality and acceptable for use in geological and resource modeling.

157

The resource modeling is well documented and the geostatistical analysis of data from the Prosperity Project Database supports the Measured and Indicated Mineral Resources listed in Table 19-1. These resources are inclusive of the stated mineral reserves.

Table 19-1

Prosperity Mineral Resources

at 0.14% Copper Cut-off

Category	Tonnes (millions)	Gold (gpt)	Copper (%)
Measured	547.1	0.46	0.27
Indicated	463.4	0.34	0.21
Total	1,010.5	0.41	0.24

Engineering Studies

The Company and its consultants carried out progressive engineering, metallurgical and environmental studies over the period 1998 to 2009 including a feasibility level study of the Prosperity Project in 2000 by Kilborn SNC Lavalin, a mill redesign and project cost review in 2006 by SNC Lavalin, a pre-feasibility study in 2007, a feasibility study with HATCH in 2007 and in 2008 Taseko worked with Axxent Engineering Ltd and Rutter Hinz to investigate value engineering opportunities.

Pre-Production and Mine Plan

The proposed project development incorporates activities during a pre-production period of two years which include construction of the electrical transmission line; upgrading and extension of current road access and mine site clearing; site infrastructure, processing, and tailings starter dam construction; removal and storage of overburden; and pre-production waste development.

The mine plan utilizes a large-scale conventional truck shovel open pit mining and milling operation. Following a one and a half-year pre-strip period, total material moved over years 1 through 24 averages 200,000 tonnes/day. A declining net smelter return cut-off is applied to the mill feed, which defers lower grade ore for later processing. The lower grade ore is recovered from stockpile for the final 2.5 years of the mine plan. The life of mine strip ratio including processing of lower grade ore is 1.5:1.

The 2009 pit optimization incorporates appropriate input parameters and the mine design incorporates an appropriate level of detail with respect to design and operating parameters.

The pit wall slopes incorporated in the current design are consistent with all geotechnical recommendations.

158

Processing and Infrastructure

The process design, recovery relationships and concentrate parameters are all supported by adequate metallurgical test work. Processing incorporates proven processes and technologies.

The Prosperity processing plant has been designed with a nominal capacity of 70,000 tonnes/day. The plant consists of a single 12-m diameter semi-autogenous grinding (SAG) mill, two 7.9-m diameter ball mills, followed by processing steps that include bulk rougher flotation, regrinding, cleaner flotation, thickening and filtering to produce a copper-gold concentrate. Expected life-of-mine metallurgical recovery is 87% for copper and 69% for gold, with annual production averaging 110 pounds copper and 234,000 ounces gold over the 33 year mine life.

The copper-gold concentrate would be hauled with highway trucks to an expanded load-out facility at McLeese Lake for rail transport to various points of sale, but mostly through the Port of North Vancouver for shipment to smelters/refineries around the world.

Power would be supplied via a new 124 km long, 230 kV transmission line from Dog Creek on the BC Hydro Grid. Infrastructure would also include the upgrade of sections of existing roads, construction of a short spur to the site, an on-site camp, equipment maintenance shop, administration office, warehouse, and explosives facilities.

The tailings storage facility is designed based on valid field data, current engineering standards, and has the capacity to contain the processed reserve and potentially acid generating materials identified in the mine plan.

The reserve estimate takes into consideration all geologic, mining, milling, and economic factors, and is stated according to Canadian standards (NI43-101).

The mineral reserves estimated from the study are as follows:

Table 19-2

Prosperity Mineral Reserves

at CDN$5.50 NSR/t Pit-Rim Cut-off

Category	Tonnes (millions)	Gold (gpt)	Copper (%)	Recoverable Gold Ounces (millions)	Recoverable Copper Pounds (billions)
Proven	481	0.46	0.26	5.0	2.4
Probable	350	0.35	0.18	2.7	1.2
Total	831	0.41	0.23	7.7	3.6

159

20. Recommendations

An Environmental Assessment Certificate is currently being pursued under the harmonized British Columbia Environmental Assessment Act (EA Act) and Canadian Environmental Assessment Act (CEAA) review process. A provincial government decision whether or not to issue the provincial Environmental Assessment Certificate will be made early in 2010 while a federal EA Certificate is not expected before mid 2010. Applications for appropriate provincial permits should be prepared so as to minimize the time between granting of a provincial EA certificate and issuance of permits.

The tailings storage facility construction design was based on the use of only non-PAG material for embankment construction. The opportunity to utilize some PAG material in sections of the embankments that remain below the phreatic surface should be investigated.

160

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Riddell, J., Schiarizza, P., Gaba, R.G.., Caira, N., And Findlay, A., 1993. Geology and Mineral Occurrences of the Mount Tatlow Map Area (920/5, 6 and 12). In Geological Fieldwork 1992. British Columbia Ministry of Energy, Mines and Petroleum Resources, Paper 1993-1, p. 37-52.

SNC Lavalin, “Prosperity Redesign and Costing Study”, August 2006

SNC Lavalin T&D, 2007, Taseko Mine Interconnection System Impact Study, completed for British Columbia Transmission Corporation

162

Talisman Land Resource Consultants INC., 1997. Prosperity Project, Preliminary Overburden and Assessment and Characterization. Unpublished Company Report, Taseko Mines Limited, Vancouver, British Columbia.

Tipper, H.W., 1969. Mesozic and Cenozic Geology of the Northeastern Part of the Mount

Waddington Map Area (92N), Coast District, British Columbia. Geological Survey of Canada,

paper 68-33.

Titley, E., 1994. Manual for Core Logging and Data Compilation on a Diamond drilling Project. Unpublished Company Report, Taseko Mines Limited, Vancouver, British Columbia.

163

22. Date and Signature Page

Signed at Vancouver, British Columbia on the 25th day of June 2010.

“Signed and Sealed”

Scott S. Jones, P.Eng.

164

Scott Jones, P.Eng.

Suite 300-905 West Pender Street

Vancouver, BC V6C 1L6

I, Scott S. Jones, P.Eng., of Vancouver, British Columbia, hereby certify that:

1. I am an employee of Taseko Mines Ltd., with a business office at 300-905 West Pender Street, Vancouver, British Columbia. In my position as Vice-President, Engineering, on behalf of Taseko Mines Limited, I co-authored this technical report on the reserve increase for the Prosperity Project which was announced on November 2, 2009.

2. This certificate applies to the technical report titled “Technical Report prepared for Franco-Nevada Corporation in respect of the Prosperity Gold/Copper Project British Columbia, Canada owned by Taseko Mines Limited”, dated June 25, 2010

3. I am a graduate of McGill University in Montreal, Quebec (B.Eng. Mining). I have practiced my profession for 24 years since graduation in 1985. I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia, license number 29486. As a result of my experience and qualifications, I am a Qualified Person under National Instrument 43-101.

4. I am responsible for the compilation of all sections of this report.

5. I am independent of Franco-Nevada Corporation.

6. I have visited the Prosperity property on multiple occasions over the period 2006 to 2009.

7. I have read National Instrument 43-101.

8. I, as of the date of the certificate and to the best of my knowledge and information, believe the technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

9. I consent to the use of his Technical report for disclosure purposes of Franco-Nevada Corporation.

Signed at Vancouver, British Columbia on the 25th day of June, 2010.

“Signed and Sealed”

Scott S. Jones, P.Eng.

Gary Giroux, PEng.

Suite 1215-675 West Hastings Street

Vancouver, BC V6B 1N2

I, G.H. Giroux, of 982 Broadview Drive, North Vancouver, British Columbia, do hereby certify that:

1) I am a consulting geological engineer with an office at #1215 - 675 West Hastings Street, Vancouver, British Columbia.

2) I am a graduate of the University of British Columbia in 1970 with a B.A. Sc. and in 1984 with a M.A. Sc., both in Geological Engineering.

3) I am a member in good standing of the Association of Professional Engineers and Geoscientists of the Province of British Columbia.

4) I have practiced my profession continuously since 1970. I have had over 30 years experience calculating mineral resources. I have previously completed resource estimations on a wide variety of porphyry deposits both in B.C. and around the world, many similar to Prosperity.

5) I have read the definition of “qualified person” set out in National Instrument 43-101 and certify that by reason of education, experience, independence and affiliation with a professional association, I meet the requirements of an Independent Qualified Person as defined in National Instrument 43-101.

6) This certificate applies to the technical report titled “Technical Report prepared for Franco-Nevada Corporation in respect of the Prosperity Gold/Copper Project British Columbia, Canada owned by Taseko Mines Limited”, dated June 25, 2010, and more specifically the resource modeling and estimation sections 17.1 and 17.2, which were prepared by others based on work I completed in Vancouver during 1998 and amended in 1999. I have not visited the property.

7) I have previously completed a resource estimation on this property in 1994.

8) As of the date of this certificate, to the best of my knowledge, information and belief, the technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

9) I am independent of Franco-Nevada Corporation applying all of the tests in section 1.4 of National Instrument 43-101.

10) I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

11) I consent to the use of this Technical report for disclosure purposes of Franco-Nevada Corporation.

Dated this 25th day of June 2010

“signed and sealed”

G. H. Giroux, P.Eng., MASc.

Lawrence A. Melis, P.Eng.

Suite 100, 2366 Ave C North

Saskatoon SK S7L 5X5

I, Lawrence A. Melis, of 259 Egnatoff Cres., Saskatoon, Saskatchewan, do hereby certify that:

1) I am a consulting process engineer, working for Melis Engineering Ltd. with an office at 2366 Ave C North, Suite 100, Saskatoon, Saskatchewan, Canada.

2) I am a graduate of the University of Western Ontario in 1971 with a B.Sc. (Chemistry).

3) I am a member in good standing of the Association of Professional Engineers and Geoscientists of the Province of British Columbia (Registration No. 19398).

4) I have practiced my profession continuously since 1971. I have had over 35 years experience in process engineering for the mining industry.

6) This certificate applies to the technical report titled “Technical Report prepared for Franco-Nevada Corporation in respect of the Prosperity Gold/Copper Project British Columbia, Canada owned by Taseko Mines Limited”, dated June 25, 2010, and more specifically a review of the metallurgy section, Section 16.0, which was prepared by others based on metallurgical testwork completed in the 1990’s by Melis Engineering Ltd. for which I was directly responsible.

6) I have visited the property in the 1990’s to look at core and general site conditions.

8) As of the date of this certificate, to the best of my knowledge, information and belief, Section 16.0 of the technical report contains all scientific and technical information that is required to be disclosed to make the metallurgical component of the technical report not misleading.

9) I am independent of Franco-Nevada Corporation as defined by National Instrument 43-101.

10) I consent to the use of this Technical report for disclosure purposes of Franco-Nevada Corporation.

Dated this 25th day of June, 2010

“signed and sealed”

Lawrence A. Melis, P.Eng.

June 25, 2010

British Columbia Securities Commission
Alberta Securities Commission
Saskatchewan Financial Services Commission
Manitoba Securities Commission
Ontario Securities Commission
Autorité des marches financiers
New Brunswick Securities Commission
Nova Scotia Securities Commission
Prince Edward Island Securities Office
Securities Commission of Newfoundland and Labrador
Government of Yukon
Government of Northwest Territories
Government of Nunavut

RE: Franco-Nevada Corporation Consent under 43-101

Pursuant to National Instrument 43-101, I, Scott Jones, P.Eng., consent to the public filing of the technical report:

Technical report prepared for Franco-Nevada Corporation in respect of the Prosperity Gold-Copper Project, British Columbia, Canada owned by Taseko Mines Limited dated June 25, 2010

I confirm that I have read the news release dated May 12, 2010 and that it fairly and accurately represents the information in the Report that supports the disclosure.

Sincerely,

Scott Jones


Scott Jones, P.Eng.

June 25, 2010

RE: Franco-Nevada Corporation Consent under 43-101

Pursuant to National Instrument 43-101, I, Gary Giroux, P.Eng., consent to the public filing of the technical report:

Technical report prepared for Franco-Nevada Corporation in respect of the Prosperity Gold-Copper Project, British Columbia, Canada owned by Taseko Mines Limited dated June 25, 2010

I confirm that I have read the news release dated May 12, 2010 and that it fairly and accurately represents the information in the Report that supports the disclosure.

Sincerely,

Gary Giroux


Gary Giroux, P.Eng.

June 25, 2010

RE: Franco-Nevada Corporation Consent under 43-101

Pursuant to National Instrument 43-101, I, Lawrence Melis, P.Eng., consent to the public filing of the technical report:

Technical report prepared for Franco-Nevada Corporation in respect of the Prosperity Gold-Copper Project, British Columbia, Canada owned by Taseko Mines Limited dated June 25, 2010.

I confirm that I have read the news release dated May 12, 2010 and that it fairly and accurately represents the information in the Report that supports the disclosure.

Sincerely,

Lawrence Melis


Lawrence Melis, P.Eng.