EXHIBIT 4

IMPORTANT NOTICE

This report was prepared exclusively for Northgate Exploration Limited (Northgate) by MRDI Canada, a Division of AMEC E&C Services Limited (MRDI). The quality of information, conclusions and estimates contained herein is consistent with the level of effort involved in MRDI’s services and is based on: i) information available at the time of preparation, ii) data supplied by outside sources and iii) the assumptions, conditions and qualifications set forth in this report. This report is intended to be used by Northgate only, subject to the terms and conditions of its contract with MRDI. Any other use of, or reliance on, this report by any third party is at that party’s sole risk.

mrdi

21 December 2001

Mr. J. Peter Gordon
Northgate Exploration Ltd.
Suite 4420 -181 Bay St
Toronto, Ontario

Dear Peter,

Re: Technical Report — Kemess Mine and Kemess North deposit

Please find enclosed our independent Qualified Person’s Technical Report on the Kemess Mine and Kemess North deposit, British Columbia. We have issued four hard copies of the report and an electronic version on CD, suitable for submittal to the relevant securities regulatory authorities.

Work on this review was done in our Toronto office. I was the Qualified Person responsible for the report preparation and conducted the independent site visit to the Kemess Mine and Kemess North deposit during November of 2001.

Yours sincerely,

Maryse Bélanger, Bsc (Geology)
Chief Geologist

MRDI Canada
111 Dunsmuir Street, Suite 400
Vancouver, B.C. V6B 5W3
Tel +1 604-664-3471
Fax +1 604-664-3041
www.amec.com

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CERTIFICATE OF AUTHOR

Maryse Bélanger
2020 Winston Park Drive, Suite 700
Oakville, ON
Tel: (905) 829-5400
Fax: (905) 829-3633
maryse.belanger@mrdi.com

I, Maryse Bélanger, am a Geologist, Chief Geologist of MRDI Canada of 337 Durie Street in the City of Toronto in the Province of Ontario.

I am a member of the Association of Geoscientists of Ontario. I graduated from the University du Quebec a Chicoutimi with a Bachelor of Science degree in geology in 1985. In 1986 I studied geostatistics at the Centre de Geostatistique de Fontainebleau for a period of 7 months.

I have practiced my profession continuously since 1987 and have been involved in: mineral exploration for gold in Canada and in geostatistics and resource modelling for copper, gold, silver, diamonds and platinum/palladium properties in Canada, United States, Philippines, Chile, Argentina, Ecuador, Russia, Ghana, Burkina Faso, Niger, Gabon, Ivory Coast, Ethiopia and South Africa.

As a result of my experience and qualifications, I am a Qualified Person as defined in N.P. 43-101.

I am currently a Consulting Geologist and have been so since March 2000.

This report was prepared under my direct supervision. I visited the Kemess Mine from May 29 to June 5, 2000 and from November 7 to 9, 2001. My most recent visit to Kemess Mine in northern British Columbia was for the purposes of reviewing pertinent geological, mining and metallurgical data in sufficient detail to independently support the Kemess Mine Mineral Reserve and Kemess North Mineral Resource statements.

This report was prepared under my direct supervision in consultation with technical specialists in the fields of mine engineering and metallurgy. Stephen Hodgson, P. Eng., President of MRDI Canada, a member of the Association of Professional Engineers and Geoscientists of British Columbia reviewed the mining aspects of this Technical Report. Mr. Hodgson graduated with a Bachelor of Science in Mineral Engineering (Mining) from the University of Alberta. He has over 25 years of experience in various aspects of mining including operations management and engineering experience at open-pit and underground mines in Canada and Alaska.

Mr. James Sorensen, President, MRDI reviewed the Kemess Mine metallurgical data. Jim has more than 20 years of experience specializing in large copper plants. A leader in technical

MRDI Canada
A division of AMEC E&C Services Limited
111 Dunsmuir Street, Suite 400
Vancouver, B.C. V6B 5W3
Tel +1 604-664-3471
Fax +1 604-664-3041
www.amec.com

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innovation of SX-EW plant operation and low-cost design, he has been responsible for the design and adaptation of column cell flotation for removal of SX organic entrainments; the development of surfactant wetting agents for copper heap leaching; full-scale onstream organic remediation through clay treatment; and continuous recovery and re-use of organic solutions from ponds. Jim has been responsible for process and design implementation on several of the world’s largest copper facilities including Lomas Bayas, Zaldivar, and Quebrada Blanca in Chile and Morenci in Arizona. He spent a number of years in operations where he was responsible for management of heap leach operations, SX-EW estimating and production forecasting, and metallurgical laboratory supervision at a 50,000 tpy copper facility.

I am not aware of any material fact or material change with respect to the subject matter of this technical report that is not reflected in this report and that the omission to disclose would make this report misleading.

I am independent of Northgate Exploration Ltd. in accordance with the application of Section 1.5 of National Instrument 43-101.

I have read National Instrument 43-101 and Form 43-101FI and this report has been prepared in compliance with same.

Dated at Toronto, Ontario, this l9th day of December, 2001.

Maryse Bélanger, Geologist
AGO Member #1491

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CONSENT OF QUALIFIED PERSON

TO:  The securities regulatory authorities of each of the provinces and territories of Canada

     I, Maryse Bélanger, Bsc (Geology), do hereby consent to the filing of the technical report prepared for Northgate Exploration Ltd. and dated December 21, 2001 in respect of the Kemess Mine and Kemess North deposit, British Columbia.

DATED at this 21st day of December 2001.

Maryse Bélanger, Bsc (Geology)

MRDI Canada
111 Dunsmuir Street, Suite 400
Vancouver, B.C. V6B 5W3
Tel +1 604-664-3471
Fax +1 604-664-3041

Northgate Exploration Limited
Kemess Mine Review
Technical Report mrdi

Table of Contents

			Page
1.0	SUMMARY		1-1
2.0	INTRODUCTION AND TERMS OF REFERENCE		2-1
	2.1	Terms of Reference	2-1
3.0	DISCLAIMER		3-1
4.0	PROPERTY DESCRIPTION AND LOCATION		4-1
5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY		5-1
6.0	HISTORY		6-1
	6.1	Kemess South	6-1
	6.2	Kemess North	6-1
7.0	GEOLOGICAL SETTING		7-1
	7.1	Kemess South	7-1
	7.2	Kemess North	7-2
8.0	DEPOSIT TYPES		8-1
9.0	MINERALIZATION		9-1
	9.1	Supergene/Leach Cap Ore	9-1
	9.2	Hypogene Ore	9-1
10.0	EXPLORATION		10-1
11.0	DRILLING		11-2
	11.1	Kemess South	11-2
	11.2	Kemess North	11-2
	11.3	Core Logging Procedures	11-6
12.0	SAMPLING METHOD AND APPROACH		12-1
13.0	SAMPLE PREPARATION, ANALYSES AND SECURITY		13-1
14.0	DATA VERIFICATION		14-2
	14.1	Kemess South	14-2
	14.2	Kemess North	14-4
15.0	ADJACENT PROPERTIES		15-1
16.0	MINERAL PROCESSING AND METALLURGICAL TESTING		16-1
	16.1	Introduction	16-1
	16.2	Background	16-1
	16.3	Metal Recovery	16-2
	16.4	Plant Availability	16-4
	16.5	Mill Throughput	16-5
	16.6	Mill Flotation Concentrate	16-6
	16.7	Conclusions	16-6
17.0	MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES		17-1
	17.1	Geologic Models	17-1
	17.2	Interpolation Plans	17-4

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			Page
	17.3	Model Validation	17-5
	17.4	Resource Classification and Summaries	17-8
18.0	OTHER DATA AND INFORMATION		18-1
	18.1	Reconciliation	18-1
19.0	REQUIREMENTS FOR TECHNICAL REPORTS ON PRODUCTION		19-1
	19.1	Mining Operations	19-1
	19.2	Process Metal Recoveries	19-2
	19.3	Markets	19-2
	19.4	Contracts	19-3
	19.5	Environmental Considerations	19-3
	19.6	Taxes	19-3
	19.7	Capital and Operating Cost Estimates	19-3
	19.8	Economic Analysis	19-4
	19.9	Payback	19-5
	19.10	Mine Life	19-5
20.0	INTERPRETATION AND CONCLUSIONS		20-1
21.0	REFERENCES		21-1
	List of Tables
	Table   1.1: 2001 Kemess Mineral Reserves and Resources		1-3
	Table   4.1: Mine Location		4-1
	Table   4.2: Kemess Mining Leases		4-1
	Table   4.3: Current and Pending Permits, Kemess Mine		4-6
	Table   9.1: Kemess Hypogene Ore Reserves Distribution by Sulphur Content and Py/Cpy Ratio		9-2
	Table 11.1: Kemess South Deposit Drilling History		11-2
	Table 11-2: Kemess North Deposit Drilling History		11-3
	Table 13.1: Kemess Analytical Analysis on Drill Core		13-1
	Table 14.1: Drill Holes Selected for 5% Database Verification		14-2
	Table 14.2: Drill Holes Selected for 5% Database Verification		14-4
	Table 16.1: Criteria for Mill Metallurgical Performance — January to October 2001		16-2
	Table 16.2: Expected Mill Operations Metallurgical Performance Criteria — March 2001		16-2
	Table 16.3: Actual Mill Metallurgical Performance Averages — 2001		16-3
	Table 16.4: Comparison of Average Hypogene Mill Metallurgical Performance Predictions		16-3
	Table 16.5: AMEC Estimated Criteria for Mill Metallurgical Performance — November 2001		16-7
	Table 17.1: Project Limits and Model Cell Block Sizes		17-1
	Table 17-2: Geology Unit Names and Block Model Codes — Kemess South		17-2
	Table 17.3: Geology Unit Names and Block Model Codes — Kemess North		17-2
	Table 17.3: Variogram Parameters — Kemess South		17-3
	Table 17.4: Variogram Parameters — Kemess North		17-3
	Table 17.5: SG Values for Selected Core Samples		17-4
	Table 17.6: Copper Mean and Standard Deviation (s.d.) Values from Ordinary Kriging and Nearest-Neighbour Estimates		17-8
	Table 17.7: 2001 Kemess Mineral Reserves and Resources		17-8
	Table 17.8: Price, Recovery and Concentrate Grade — Kemess South		17-9
	Table 18.1: Summary of F1, F2 and F3 Factors for Kemess South, 1998 to 2000		18-2
	Table 18.2: Year-End 2000 Reconciliation (provided by Kemess)		18-3

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	Page
Table 18.3: Year 2001 Copper Reconciliation Data	18-3
Table 18.4: Year 2001 Gold Reconciliation Data	18-4
Table 19.1: Production Forecast	19-2
Table 19.2: AMEC Estimated Criteria for Mill Metallurgical Performance — November 2001	19-2
Table 19.3: Projected Operating Costs	19-4
List of Figures
Figure   4.1: Kemess District Location Map	4-2
Figure   4.2: Deposit Location Map	4-3
Figure   4-3: Claim Map	4-4
Figure   4.4: Air Photo Infrastructure Map	4-5
Figure   7.1: Pit Plan at 1230 Elevation	7-3
Figure   7.2: Pit Longitudinal Section at 9900N	7-4
Figure   7.3: Pit Cross Section at 10000E	7-5
Figure   7.4: Longitudinal Section for Kemess North — Genetic model	7-6
Figure   7.5: Cross-Section for Kemess North — 2001 Drilling Program	7-7
Figure 11.1: Drill Hole Locations — Kemess South	11-4
Figure 11.2: Drill Hole Locations — Kemess North	11-5
Figure 17.1: Grade-Tonnage Curves for Copper	17-6
Figure 17.2: Grade-Tonnage Curves for Gold	17-7
Appendices
Appendix A                               Assay Procedures
Appendix B                               Sections

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1.0	SUMMARY
	Northgate Exploration Limited (Northgate, or the Company) has asked MRDI Canada, a division of AMEC E&C Services Limited (MRDI), to provide an independent Qualified Person’s Review and Technical Report of the Kemess Mine in British Columbia. Maryse Bélanger, who is a member of the Association of Geoscientists of Ontario (AGO) and an employee of MRDI, served as the Qualified Person responsible for the preparation of the Technical Report as defined in National Instrument 43-101, Standards of Disclosure for Mineral Projects, and in compliance with Form 43-101F1 (the “Technical Reports”). The work entailed review of pertinent geological, mining and metallurgical data in sufficient detail to prepare the Technical Report, the purpose of which is to independently support the 2001 mineral resource and reserve statement.
	At this time, two deposits are defined at Kemess: Kemess South, which is currently under production as an open pit mine, and Kemess North, an undeveloped exploration project. The Mineral Reserves are contained exclusively in the Kemess South deposit. Kemess North comprises Inferred Mineral Resources. Production from the Kemess South deposit began in 1998, and it continues to be the sole location of mining.
	Information and data for the review and report were obtained from the Kemess Mine during two site visits by MRDI, one from 29 May to 5 June 2000, and another from 7 to 9 November 2001. Background information for property description, land status, infrastructure, property history, geology and mineralization was obtained from existing company reports, papers and drawings and discussions with Northgate personnel at the mine site. Detailed data on resource and reserve matters (database, resource modelling, production reconciliation, mine design, production costs, metallurgical data, process costs and general and administrative costs) were made available and examined during the site visits. Additional information concerning economic factors in support of the mineral reserve designation was obtained through discussions with Northgate senior management in Toronto, Ontario.
	Kemess South and Kemess North are typical porphyry gold-copper deposits associated with the presence of monzonitic to monzodioritic intrusions in the surrounding Takla volcanics. The mineralization in these deposits is contained in stockworks, veins, breccias and disseminations of pyrite, chalcopyrite, bornite and magnetite that are intimately associated with hydrothermal alteration of the intrusive body and the volcanic country rocks.
	The database used to estimate the Mineral Reserves consists of samples from 223 core holes. Samples were assayed for copper and gold. Data transfer from the

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original assay certificates to the resource database was verified through a random check of 5% of the data points. Original copper analyses are supported by production history (reconciliation of the block model to blast holes and to milled ore). In 2001, there was a 14% positive variance in gold head grades with the block model, a phenomenon not previously observed when mining the upper benches.

An average specific gravity was developed for each rock type in the Kemess South deposit based on a series of measurements performed by the water displacement method (wax immersion method). Reconciliation data show that the Kemess South block model provides an accurate estimate of reserve tonnes.

The mineral resource and reserve estimates for the two deposits at the Kemess Mine were calculated under the direction of Michael Hibbitts, Northgate’s Qualified Person for mineral resources and reserves. The estimates were made from 3-dimensional block models utilizing commercial mine planning software (MineSight®). Industry-accepted modelling procedures were used to create interpolation domains based on mineralized geology and grade estimation based on ordinary kriging. Reasonableness of grade interpolation was reviewed by visual inspection of sections and plans displaying block model grades, drill hole composites and geology. Good agreement was observed. Model checks on smoothing were also performed, and an appropriate amount of internal dilution was found to be included in the mineral reserve.

The block model results were compared to blast hole and milled production data. Blast hole drill data for year 2001 production were compared to the block model for the same areas: good agreement for average grade copper values was observed. Gold production exceeded expectations by an average of 15%. Annual reconciliation and comparison of milled production (mill measured grades and tonnage) to the block model (calculated from year-end topography) also shows good agreement for copper.

Classification of the mineral resource and reserve at Kemess is based on a combination of drill data density, economic evaluation and production. The Kemess South deposit contains an appropriate density of drilling (100 m x 100 m), good copper validation results, especially to blast holes and milled production, and well-known economic and metallurgical performance (through mining history). This supports the category of Probable Mineral Reserve for all Kemess South deposit mineralization above cutoff within the ultimate pit outline.

The Kemess North deposit is appropriately classified as an Inferred Resource, reflecting the lower drill hole density and resultant higher uncertainty with respect to grade and geologic continuity. Limited metallurgical testwork has been undertaken for

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Kemess North (see Section 16), but no economic evaluation has been completed as yet.

The total Kemess mineral reserves and resources as of 31 December 2001 are summarized in Table 1.1.

Table 1.1:  2001 Kemess Mineral Reserves and Resources

					Grades				Contained Metals
	Cutoff Grade
	Gold (gmt)				Gold		Copper		Gold		Copper
	Equivalent		Tonnes		(gmt)		(%)		(oz)		(M lb)
Inferred Mineral Resource
Kemess North		0.6		442,000,000		0.40		0.23		5,700,000		2,200
		0.8		170,000,000		0.50		0.29		2,700,000		1,100
Mineral Reserve
Kemess South				145,911,000		0.65		0.24		3,063,000		756

The gold equivalent calculations for Kemess North are based on metal prices of U.S.$325 per ounce gold and U.S.$0.90 per pound copper.

The long-term production forecast is based on a mill throughput of 17.5 million tonnes of ore per year and the production of approximately 155,000 tonnes of copper concentrate. The current life-of-mine plan is based only on mineral reserves and forecasts that the Kemess South deposit will be mined out in 2009.

Operating costs have been forecast as a function of previous costs. Details of economic and sensitivity analyses, marketing and concentrate sales contracts, taxes, environmental and geotechnical considerations were reviewed. All were complete and fall within accepted industry standards.

The 2001 Kemess Mineral Reserve and Mineral Resource statement is supported by this independent review.

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2.0	INTRODUCTION AND TERMS OF REFERENCE
	Northgate Exploration Limited (Northgate) commissioned MRDI Canada, a division of AMEC E&C Services Limited (MRDI), to provide an independent Qualified Person’s Review and Technical Report of the Kemess Operation. Maryse Belanger, an employee of MRDI, served as the Qualified Person responsible for the preparation of the Technical Report as defined in National Instrument 43-101, Standards of Disclosure for Mineral Projects, and in compliance with Form 43-101F1 (the “Technical Reports”).
	Information and data for the review and report were obtained from Kemess staff during a visit to site by MRDI on November 7 to 9, 2001. Additional information was obtained from Northgate’s office in Toronto, Ontario.
	The work entailed review of pertinent geological, mining and metallurgical data in sufficient detail to prepare the Technical Report. Maryse Belanger, in addition to supervising the preparation of the Technical Report, conducted and supervised the review of the geological data. Qualified Person assistance was provided by MRDI employees Stephen Hodgson, P. Eng., who supervised the investigation of mining issues, and James Sorensen, who conducted the review on metallurgical matters.
2.1	Terms of Reference
	The Kemess South deposit is currently being mined. Therefore, Kemess South and Kemess Mine may be used interchangeably in this report. Kemess North represents a separate, unexploited deposit, which is still undergoing delineation.
	The S.I. system of measurement is used at the Kemess Mine.

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3.0

DISCLAIMER

MRDI’s review of the Kemess operation relied on geotechnical work carried out by geotechnical experts from Knight Piesold in Vancouver. The tailings dam design and monitoring is being performed by AMEC Earth & Environmental, also of Vancouver. MRDI used information from these firms under the assumption that it was prepared by Qualified Persons. The reports prepared by Knight Piesold and AMEC Earth & Environmental are listed in Section 21.

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4.0	PROPERTY DESCRIPTION AND LOCATION
	The Kemess Mine is 430 km northwest of Prince George, British Columbia approximately 57°02’ north longitude, 126°47’ west latitude on NTS map 94/e2 (Table 4.1 and Figure 4.1). It comprises two deposits, Kemess South and Kemess North, which are 5.5 km apart (Figure 4.2).
	The elevation at the Kemess Mine property generally varies from 1,300 m to 1,800 m above sea level. The highest point in the region is a westerly trending glacial cirque at 2,000 m elevation. The Kemess South deposit is below the local tree line. The Kemess North deposit is in an alpine environment approximately 100 m to 150 m above the tree line.
Table 4.1: Mine Location

Deposit	Longitude		Latitude
Kemess South		57°02’		126°47’

The Kemess property encompasses 77,000 acres (31,160 hectares) included in various claims listed in Table 4.2. The mine is covered under Mining Lease 354991, which is active until September 2015. The Kemess Mine leases are shown in Figure 4.3.

Table 4.2: Kemess Mining Leases

Leases	Tenure #		Tag #		Map #		Expiry Date	Units		Hectares		Acres
New Kemess No. 1		237800		9355		094E02W	15-Dec-10		18		450		1111.97
New Kemess No. 2		237801		9356		094E02W	15-Dec-10		20		500		1235.52
New Kemess No. 3		241960		120227		094E02E	15-Dec-10		15		375		926.64
Mining Lease for Kemess South		354991		—		094E02E	Sept 2015		863		—		—

Infrastructure at the Kemess Mine consists of an office and maintenance building, a mill building, access and service roads and an airstrip. These are shown in Figure 4.4 along with the locations of the open pit, waste rock dumps and tailings impoundment area.

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Figure 4.1: Kemess District Location Map

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Figure 4.2: Deposits Location Map

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Figure 4.3: Claim Map

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Figure 4.4: Air Photo Infrastructure Map

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Numerous permits necessary for the operation of mining activities at the Kemess Mine have been either granted or pending. No outstanding permits need to be obtained. The permits in place or pending are listed in Table 4.3.

Table 4.3: Current and Pending Permits, Kemess Mine

Permit	Title	Date of Issue
M96-03	Project Approval Certificate	29 April 1996
M206	Approving Work System &	July 1996, amended
	Reclamation Program	21 January 2000
PE15335	Main Effluent Permit -- Tailings	8 December 1998,
	Storage Facility & Associated	amended 20 July 2000
	Works, RBC, Mill and
	Accommodation Site Runoff &
	Open Pit Water
M206 Amendment	Waste Rock Dump Permit	Pending
Section 35-2	Fisheries (DFO) Authorization	4 November 1996
BCG07761	Special Waste Consignor	18 February 1998
	Identification Number
Explosives Storage and	Main magazine storage of	21 January 1998
Use Permits n* 682	explosives and detonators
Explosives Storage and	Magazine storage for avalanche	14 February 1997
Use Permits n* 1168	explosives and detonators
	Air Emissions	Pending
	Land Farming	Pending
01-7829.01-99	Finlay Road	4 February 1999

A discussion of environmental liabilities and management procedures for the Kemess Mine is included in Section 19.5 of this report under the subheading “Environmental Considerations.”

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5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY
	The Kemess mining and milling complex is located in the mountains of north-central British Columbia at an elevation of 1,350 m. Personnel access to the mine is by plane via Prince George and Smithers, with flight service available from Monday through Thursday. Road access to the mine is from Mackenzie, B.C., and is the means by which supplies and concentrates are hauled to and from the mine. The Kemess North deposit is accessed by helicopter from the mining and milling complex to the south. Power at the site is available directly from the BC Hydro grid, the British Columbia power authority, via a Company-owned 380 km powerline.
	The Kemess Mine belongs to the following physiographic subdivision of British Columbia, arranged in ascending hierarchy of units: Swannell Ranges, Omineca Mountains, Central Plateau and Mountain Area, Interior System and Canadian Cordillera.
	Two biogeoclimatic zones occur in the Kemess Mine area, according to biogeoclimatic maps of the Toodoggone River 94E and McConnell Creek 94D map sheets. The mild, cool Spruce-Willow-Birch Zone (SWB mk) occupies the lower elevations between 1,200 m and 1,500 m. Most of the mine is within this zone. The Alpine Tundra parkland (Atp) subzone occupies the higher elevations in the mine area.

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6.0	HISTORY
6.1	Kemess South
	Pacific Ridge Resources Ltd. (Pacific Ridge) staked the area of the Kemess South deposit in 1983. Exploration programs were subsequently carried out by Pacific Ridge Resources Ltd. and Anaconda Canada Ltd. (Anaconda) in 1984; St. Philips Resources Inc. (St. Philips) in 1988 and the Kemess South Joint Venture between El Condor Resources Ltd. (El Condor) (60%), operator, and St. Philips Resources Ltd. (40%) from 1990 to 1993. In 1991, Rio Algom Explorations Inc. (Rio Algom) acquired claims adjoining the west and south sides of the Kemess South Joint Venture claim holdings.
	The initial work on the property by Pacific Ridge and Anaconda consisted of a limited diamond-drilling program to test a gold-copper-molybdenum soil geochemical anomaly. This drilling identified porphyry-style gold-copper-molybdenum mineralization, but grades were considered too low and the property was dropped. St. Philips carried out IP surveys, geochemical surveys and reverse circulation drilling, which marginally expanded the mineralized area. The Kemess South Joint Venture completed a major delineation diamond drilling program and various ancillary works, including IP and geochemical surveys. In 1992, Rio Algom drilled five holes totalling 1,745 m to further delineate the deeply buried western extension of the Kemess South deposit. In late 1993 the Kemess South Joint Venture acquired the claims held by Rio Algom. By the end of 1993 a total of 26,314 m of diamond drilling in 156 holes had outlined a substantial gold-copper deposit that was amenable to open pit development.
	In 1994 the Kemess South Joint Venture conducted a 9-hole, 1,867 m in-filling drilling program. In 1996, Royal Oak Mines Inc. (Royal Oak) acquired the Kemess South property and drilled 22 due diligence holes totalling 3,316 m.
6.2	Kemess North
	The Kemess North deposit was discovered prior to Kemess South when Kennoco Exploration (Western) Limited (Kennoco) began exploring in the region during 1966. Kennoco carried out a regional silt geochemical survey in the vicinity of the New Kemess 2 claim. The following year, Kennoco staked 100 two-post mineral claims to cover an intense gossan with high base metal and silver silt geochemistry.
	From 1968 to 1971 Kennoco conducted a more detailed soil, silt and rock geochemical sampling program in the area. A program of geological mapping program was completed at a scale of 1:9,600. The company also drilled eight X-ray diamond drill holes (232 m), but no drill core assay data were reported.

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Between 1975 and 1976, Getty Mines Limited and Shell Oil jointly optioned the property from Kennoco. These companies restaked some of the claims that had lapsed in the area and completed a photogrammetric topographic map at 1:4,800 scale, as well as an in-fill soil geochemical sampling and geological mapping program. Diamond drilling was also done, totalling 2,065 m in 13 drill holes and 192 reported assays in 1975, and 1,477 m in 7 drill holes and 491 reported assays in 1976.

From 1986 to 1992, El Condor optioned the claim group from Kennoco and carried out diamond drilling, geophysics and various mapping and sampling programs. In 1986, El Condor completed a magnetics survey covering 14.1 km, analysis of 351 soil samples and the relogging and resampling of the 1975 and 1976 drill core. Work in 1987 consisted of a rock sample program that gathered 345 rock chip samples. During 1988 El Condor completed 50.5 km of Em-34 resistivity surveying on three grids, 1.8 km of IP surveying and an extensive soil sample program covering 1,675 soil samples and 37 rock chip samples on the project grid. Ninety metres of hand trenching were also mapped and sampled. The comprehensive work programs continued in 1989, and a total of 385 m of backhoe trenching was done and sampled (246 samples). Soil geochemistry work continued, with 828 soil samples collected. Geophysical surveys during this period consisted of 26.35 line kilometres of VLF EM and magnetic surveying and 11.8 line kilometres of IP surveying. Five drill holes totalling 732 m and providing 327 assays were completed.

In 1990 El Condor conducted a 1:10,000 scale reconnaissance geological mapping program and continued with its geochemical sampling. For geochemical and heavy mineral analyses, 239 rock chip samples were collected on a 200 m x 200 m pattern. The company also drilled 12 diamond drill holes for a total of 2,207 m and analyzed 1,071 drill core assays. Additional claims were staked and optioned to cover the entire prospective area lying between the Kemess North and Kemess South deposits and several prominent gossans to the east. El Condor continued this program in 1991, drilling 21 NQ/HQ diamond drill holes totalling 6,165 m into the Kemess North deposit and 6 NQ diamond drill holes totalling 900 m into the Kemess Nugget Zone. A total of 3,370 drill core samples were analyzed. El Condor also cut 167 km of grid lines and completed 150 km of geological mapping and 147.8 km of IP surveys. In 1992 the company drilled 4,250 m in 25 NQ drill holes and assayed 2,240 samples.

In 2000, Northgate completed 12 NQ/HQ holes for a total of 4,100 m and assayed 1,958 samples. Further drilling in 2001 included 16 holes for a total of 8200 m.

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7.0	GEOLOGICAL SETTING
	The Kemess Mine is located in a northwest-trending geological belt, known as the Quesnel Trough, which extends over a distance of approximately 1,200 km in British Columbia. This feature contains several gold-copper and copper-molybdenum deposits, including the Similco and Brenda mines in the south and Mt. Polley and Mt. Milligan in the north.
	The main rock units in the Kemess District are the upper Triassic to Lower Jurassic Takla Group. These rocks are predominantly subaqueous volcanic strata consisting of lava flows with subordinate interbeds of tuff and volcanic breccia. Overlying the Takla Group are Lower Jurassic-Hazelton Group-Toodoggone Formation volcanic rocks dominated by flows and tuffs. Numerous stocks and dykes of Lower to Mid-Jurassic age intrude the Toodoggone and Takla strata.
	On the Kemess properties, the overlying Toodoggone volcanic rocks have been removed by erosion, exposing several large monzonite intrusions with disseminated sulphide mineralization and associated hydrothermal alteration. The resulting disseminated sulphide system measures at least 9 km north-south and 5 km east-west. It contains both the Kemess South and Kemess North deposits.
7.1	Kemess South
	Two ore types, a supergene zone and a hypogene zone characterize the Kemess South deposit. The supergene makes up approximately 20% of the orebody and formed during a period of extensive weathering of the underlying hypogene, which represents the 80% balance of the deposit. Overlying both zones is a highly leached cap.
	Disseminated grains of chalcopyrite, bornite and chalcocite characterize the hypogene ore. Gold is intimately associated with the copper-bearing sulphides as free, fine grains of electrum (gold/silver) and native gold. Characteristically, the gold is very fine grained, 5 µm to 50 µm, and is readily recoverable by standard copper flotation processes.
	The supergene ore is characterized by strong hematite alteration from the oxidation of the sulphide mineralogy (chalcopyrite, bornite and pyrite). The supergene ore is composed of a typical suite of supergene minerals — native copper, chalcocite, bornite, and minor accessory malachite, cuprite and azurite. Gold is found as fine native grains of gold and electrum and as copper/gold alloys (auricuprides).

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	Plan and sections prepared by the Kemess geology staff show the general geology of the Kemess South deposit. These are reproduced as Figure 7.1 (plan at elevation 1,230 m), Figure 7.2 (longitudinal section at 9900N) and Figure 7.3 (cross-section at 10000E).
7.2	Kemess North
	A large porphyry system also exists at Kemess North. The geology is similar to that of the Kemess South deposit but most probably lies 200 m to 300 m lower in the regional stratigraphy. The mineralization outlined to date is hypogene in nature and surrounded by a large-scale pyritic halo. The porphyry material is monzonitic/ monzodioritic in composition. A higher-grade core has been identified and is associated with the presence of siliceous sulphide breccia. The different alteration phases present consist of silicification with local silica flooding, clay, magnetite and sericite alteration. In some cases, the silica obliterates the siliceous sulphide breccia, and no original texture remains. Mineralization includes pyrite, chalcopyrite, magnetite, and trace bornite and native copper. Northgate has commissioned a petrographic study that will provide details of the mineralogy and alteration minerals. Representative longitudinal and cross-sections through Kemess North are provided in Figures 7.4 and 7.5.

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Figure 7.1: Pit Plan at 1230 Elevation

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Figure 7.2: Pit Longitudinal Section at 9900N

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Figure 7.3: Pit Cross-Section at 10000E

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Figure 7.4: Longitudinal Section for Kemess North — Genetic Model

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Figure 7.5: Cross-Section for Kemess North — 2001 Drilling Program

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8.0	DEPOSIT TYPES
	Kemess South and Kemess North are typical porphyry gold-copper deposits. These deposits are generally hosted in or near intrusive rocks, usually of tonalitic or quartz monzonitic composition with associated volcanic clastics and flows, and are often large, oval and inverse shaped. Typical dimensions are in the order of 1,000 m x 1,000 m x 100 m.
	The mineralization in these deposits is normally associated with stockworks, veins and disseminations of pyrite, chalcopyrite, bornite and magnetite that are intimately associated with hydrothermal alteration of the intrusive body and the volcanic country rocks. Copper porphyry deposits can display multiple zones of hydrothermal alteration and sulphide mineralization. The hydrothermal alteration is usually extensive and can consist of biotite, potassium-feldspar, sericite, anhydrite/gypsum, magnetite, hematite, actinolite, chlorite, epidote and carbonate (Panteleyev, 1995).
	Chalcopyrite, bornite, magnetite, pyrite, gold and silver are typically the dominant sulphides and metals. The mineralization is dominantly structurally controlled, mainly through stockworks, veins, vein sets, breccias, disseminations and replacements.

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9.0	MINERALIZATION
	The Kemess South mineralization is well understood, and mining has provided some valuable information on the ore characteristics. A supergene/leach cap zone overlying the hypogene material was identified during early exploration drilling.
	Only hypogene mineralization has been discovered to date at Kemess North. More work will be necessary to fully characterize it.
9.1	Supergene/Leach Cap Ore
	The supergene and leach cap ores are reported as being mineralogically similar. These materials make up only 10% to 15% of the remaining material to be mined, tend to be mined together and campaigned as combined feed to the mill. Native copper and copper-gold alloys (auricuprides) are the main copper mineral species, with minor chalcocite and trace chalcopyrite. Gold is mainly present as copper-gold alloys and free gold or electrum.
9.2	Hypogene Ore
	Representing 80% to 90% of the remaining material to be mined at Kemess South, the hypogene zones can be spatially characterized by their relative sulphur/sulphide (mainly pyrite) content. When sulphur/sulphide contents are lower, metallurgical control techniques for pyrite can have a negative influence on recovery, as well as wasting lime (overusing lime and depressing copper and gold values that would ordinarily be recoverable).
	High pyrite content materials, roughly characterized as >2% sulphur, are usually found in the northwest sector of the Kemess South deposit. This material is generally lower in grade and comprises only a quarter of the remaining reserve. The high pyrite material is visually distinguishable from the lower sulphide zones. Most of the hypogene material is <2% sulphur. The daily reports prepared by geologists at Kemess indicate the relative sulphur/sulphide content, pyrite to chalcopyrite (Py/Cpy) ratio and material type for the material to be mined each day. The hypogene ore can be broken down into the categories shown in Table 9.1.

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Table 9.1: Kemess Hypogene Ore Reserves Distribution by Sulphur Content and Py/Cpy Ratio

Category of Resource	M		% of Total		Grade		Grade		Grade		Py/Cpy
(% Sulphur)	Tonnes		Tonnes		(% Cu)		(g/t Au)		(% S)		Ratio
< 1.0% Sulphur		26.1		23		0.256		0.685		0.74		1.22
1.0 to 1.5% Sulphur		39.0		35		0.258		0.720		1.24		2.46
1.5 to 2.0% Sulphur		24.1		21		0.214		0.570		1.72		4.55
> 2.0% Sulphur		23.9		21		0.181		0.487		2.48		8.42
Total		113.1		100		0.232		0.631		1.49		3.88

Chalcopyrite is the main copper mineral species, with minor bornite. Gold is present mainly as free gold or electrum.

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10.0	EXPLORATION
	Exploration work in the Kemess district started in 1966 and has been nearly continuous, although the property ownership changed a number of times. Exploration has included geological mapping, geochemistry, geophysical surveys and drilling.
	Geophysical methods included airborne EM, resistivity, VLF and IP. The original IP surveys were generally small surveys of specific claims or geological targets.

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11.0	DRILLING
	Since 1966, the various property owners have drilled 66,304 m in 334 holes in the district (see Tables 11.1 and 11.2). This drilling consists of 62,988 m in 322 drill holes for exploration and 3,316 m in 22 drill holes for deposit delineation. A map showing the collar locations of drill holes in the Kemess South and Kemess North deposits is provided in Appendix C.
	On completion of drilling, the exploration holes at Kemess North are surveyed by the mine surveyors. All 2000 and 2001 exploration holes on the Kemess North property have been surveyed. These holes have also been down-hole surveyed at 50 m intervals and at the bottom of the hole using a single-shot Sperry Sun instrument. The most recent holes are vertical or close to vertical, with dips ranging from 75° to 90°. The only drill holes that were down-hole surveyed prior to 2000 were those from the 1991 campaign.
11.1	Kemess South
	Since 1984, 37,103 m in 223 diamond drill holes have been completed at the Kemess South deposit (Table 11.1). Figure 11.1 is a location map showing drill hole locations relative to the pit location at the Kemess Mine. The drill spacing in the pit is nominally 100 m x 100 m on 50 m spaced sections.

               Table 11.1: Kemess South Deposit Drilling History

Year	1984		1988		1990		1991		1992		1994		1996		Total
DDH		5		16		23		111		37		9		22		223
Metres		315		1,185		3,867		19,406		7,146		1,868		3,316		37,103

11.2	Kemess North
	Since 1975, 111 core holes for 29,196 m of core have been drilled in Kemess North (Table 11.2). Drill spacing is nominally 100 m x 100 m. Figure 11.2 shows the locations of the drill collars relative to the boundaries of the mineralization.

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               Table 11-2: Kemess North Deposit Drilling History

Year	1975		1976		1989		1990		1991		1992		2000		2001		Total
DDH		13		7		5		12		21		25		12		16		111
Metres		2,065		1,477		732		2,207		6,165		4,250		4,100		8,200		29,196

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Figure 11.1: Drill Hole Locations — Kemess South

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Figure 11.2: Drill Hole Locations — Kemess North

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11.3	Core Logging Procedures
	Core is delivered to the core logging building at the Kemess Mine complex, where a geologist completes a written log for the hole that includes geological and geotechnical information. All drilling, logging and sampling at the Kemess Mine is in the S.I. system of measurement.
	The geological data include identification of specific geological formations, colour, presence and visual estimate of sulphide minerals, nature of fracture filling and a detailed geological description of the core that includes textural and lithologic characteristics, deformation and contact styles, and mineralization.
	Geotechnical data are also recorded. These data include recovery, rock quality designation (RQD), joint condition, weathering and hardness. Core recovery is calculated by cumulatively measuring the pieces of core between two footage markers in the core box. This measured length is then described as a percentage of the actual core length as indicated by the footage markers. RQD measurements are captured for pieces of core that are greater than twice the core diameter. For HQ drilling the core pieces must be greater than 7.6 cm long and for NQ drilling, 10.2 cm. RQD is calculated as a percentage of all core pieces greater in length than the minimum requirement for each 2 m length of the drill hole. Fracture intensity is measured by counting the number of open fracturing (actual breaks in core competency) for every 20 cm of core.
	Specific gravity is measured at varying intervals in mineralized sections. A 20 cm long sample is selected, measured and weighed using the wax immersion method to provide a specific gravity calculation.

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12.0

SAMPLING METHOD AND APPROACH

Various sampling protocols were used for the work at Kemess North. Holes drilled in 1975 and 1976 were sampled in 3 m intervals. In 1989 samples were collected in 1.5 m intervals. In the more recent work in 2000 and 2001, samples were taken systematically at 2 m spacing. Northgate geologists sample the full length of core drilled. Standards and duplicates are inserted by the assay labs. Duplicate analyses are done at independent labs, including pulp duplicates of selected samples.

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13.0	SAMPLE PREPARATION, ANALYSES AND SECURITY
	The samples processed at the assay laboratory at the Kemess Mine are dried at 105°C for 3 hours and then crushed to minus 10 mesh using a Rhino jaw crusher. Crushed samples are riffled to approximately 200 g using a 1/4” Jones Riffle. The remaining portions of the crushed samples are placed in clean, labelled plastic bags for storage. The riffled 200 g samples are pulverized to 90% passing minus 150 mesh to make a sample pulp. The prepared pulps are packed in 5 gallon plastic pails, the covers sealed with tape and the sealed pails then shipped by bonded air courier to an independent commercial lab in Vancouver for gold and copper analyses. Copper and gold assays for the 2000 program were done by Assayers Canada Laboratory in Vancouver and for the 2001 program by Bondar-Clegg Laboratory in North Vancouver. All samples are assayed for gold by fire assay and atomic absorption techniques and for copper by atomic absorption. Detailed assaying procedures are included in Appendix A.
	The Kemess lab performs numerical control checks when the drill core samples are received for sample preparation and sample pulp packing. Both Assayers Canada and Bondar-Clegg also conduct control checks upon receipt of the pulps. Assay certificates produced by the commercial labs were cross-checked by MRDI with the sample submission sheets used by Kemess geologists.
	Table 13.1: Kemess Analytical Analysis on Drill Core

Element	Analytical Method	Detection Limit
Copper	Bromide/AAS	0.001	%
Gold	Fire assay/AA	0.01	g/t

All coarse rejects are stored inside at the mine site.

MRDI reviewed the QA/QC data from the 2000 to 2001 drill programs at the Kemess North deposit and found the results for copper to be well within acceptable limits. The gold grades show greater variability, but no systematic bias was detected. It is MRDI’s opinion that more sampling and assaying work will be necessary to upgrade the Kemess North Inferred Resources to an Indicated Resource.

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14.0	DATA VERIFICATION
	As a test of data integrity, the data used to estimate the Kemess resource and reserves were checked. MRDI checked 5% of the drill hole data in the Kemess database.
14.1	Kemeses South
	The Kemess South drill hole assay database contains 223 drill holes. Of this set, 12 holes were chosen at random to represent 5% of the total database. The 12 holes were selected from a subset of approximately 160 holes covering the Kemess South open pit area. The check holes represent drilling from 1991 through 1996 and are equally spaced across the pit. The check drill holes are listed in Table 14.1.
	Table 14.1: Drill Holes Selected for 5% Database Verification

Drill Hole	Easting		Northing
91-40		10400.9	9950.2
91-57		10549.6	9700.0
91-73		10050.9	9899.0
91-116		9653.5	1333.4
91-120		9748.6	9794.6
91-130		9451.4	9803.3
91-134		9850.1	10052.2
94-183		9898.9	9948.1
96-016		10099.6	9649.8
96-019		10302.9	9751.1
RlO92-17		9349.2	9705.4
RlO92-19		9349.3	9809.1

The database used to construct the current geological model was checked against the original core description logs and laboratory assay certificates. The following database items were scrutinized for transcription errors, missing data or additional data with no original documentation:

o Copper assay value (CUORG) on % basis

o Native copper assay value (CUNAT) on % basis

o Copper oxide assay value (CUOX) on % basis

o Gold assay value (AUORG) in grams per tonne

o Geologic domain (DOMN) including overburden, tertiary sediments, takla sediments, takla volcanics, leach cap ore, supergene ore and hypogene ore.

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The results of the 5% database checks showed minimal errors. More than 1,700 analyses were checked for the primary resource assays of gold (AUORG) and copper (CUORG). Exclusive of Rio Algom holes, only two data points were not confirmed because the data were missing on the original lab assay certificate. No original lab assay certificate could be located for drill hole 94-183, and it was checked from sample summary forms; no errors were found in this hole.

Two holes drilled in 1992 by Rio Algom were included as part of the 5% checks, although complete historical documentation was lacking for these holes. Of 380 gold and copper values, only 41 analyses could be confirmed. Rio Algom drill holes represent 5 of 160 holes within the Kemess South deposit, all 5 holes located on the western edge of the ultimate pit, outside of the defined mineable reserve. MRDI believes that the two holes included in the 5% checks have little, if any, effect on the accuracy of the resource model.

The check of secondary resource assays of native copper (CUNAT) and oxidized copper (CUOX) revealed no errors in holes prior to 1996. The database contains no CUNAT or CUOX lab assay data for any of the 1996 holes in the Kemess South area, including holes 96-016 and 96-019, even though lab assays were completed for these two holes. Although CUNAT and CUOX are not currently part of the geological model, MRDI recommends that all available assay values for native and oxide copper from the 1996 drilling program be entered in the database for future model updates. No CUNAT or CUOX assay values have been determined for the two Rio Algom holes; there is no documentation to confirm this.

The geologic domain (DOMN) codes were checked using the original drill core logging forms. Minor variations in contact depths between the originally logged domains and the present database were noted for a number of drill holes. These changes appear to be reasonable and are documented in the original drill hole file.

Drill hole collar survey data in the database were checked against the original survey certificates where possible. Field verification could not be carried out because the original topographic surface has been modified by pit development. Variations in elevation of approximately 10 m were noted between the original survey and the database for 1991 holes. This discrepancy was clarified in documents provided by El Condor Resources, which converted the survey data for older holes to the universal coordinate system. All the drill holes within Kemess South deposit were plotted on the original topographic surface as an additional check. Only two drill hole collars were found to have elevation differences of approximately 4 m when compared to the topographic surface.

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14.2	Kemess North
	The Kemess North exploration assay database contains 111 drill holes. Of this set, 5 holes were selected from the year 2000 and 2001 drilling campaigns to represent approximately 5% of the total database. The check drill holes are listed in Table 14.2.
	Table 14.2: Drill Holes Selected for 5% Database Verification

Drill Hole	Easting		Northing
KN-01-03		10265.8		16117.5
KN-01-09		10466.6		16367.2
KN-01-10		10581.5		16289.3
KN-01-12		10259.6		16277.4
KN-00-05		10162.2		15963.8

The database used to construct the current geological model was checked against the original laboratory assay certificates. The following database items were scrutinized for transcription errors, missing data or additional data with no original documentation:

o Copper assay value (CUORG) on % basis

o Gold assay value (AUORG) in grams per tonne

No errors were found when checked against the original assay certificates in the drill logs.

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15.0	ADJACENT PROPERTIES
	Adjacent properties are not relevant for this review of the Kemess South and Kemess North deposits.

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16.0	MINERAL PROCESSING AND METALLURGICAL TESTING
16.1	Introduction
	This section of the Technical Report addresses the Kemess milling facilities and those aspects of the mine site operations directly related to metallurgical performance of the mill. The review was based on various documents provided by Northgate managers and on observations made during site visits in May and August 2000. Two separate reviews were also performed, one in March 2001 and an update in December 2001. A complete list of documentation collected and used for this report is provided in Appendix C.
16.2	Background
	MRDI completed an initial draft report in June 2000. An updated report was prepared in March 2001, presenting findings and opinions based on additional operating information provided by Northgate, direct observations made during a site visit (August 2000) and subsequent support data obtained from Northgate for milling operations through December 2000. The March 2001 report represented the final document in a series of successive revisions produced to include new data as they became available over the course of fall (2000) and early winter (2001).
	The March 2001 review of the pit geology, mine plans and data indicated that most of the remaining material to be mined and processed beyond 2000 was hypogene ore. At the time, the mine had planned to deliver hypogene material to the primary crusher for approximately 11 months each year. This expectation was not realized in 2001, as feed to the mill was reportedly hypogene 49.7% of the time (166 of 334 days) and combined transition (26.9%) and supergene/leach cap (23.4%) materials the rest of the time. The supergene and leach cap (very little leach cap remains) are generally mined together and therefore combined as one material type for review purposes.
	The non-hypogene feed was processed primarily from mid-April to mid-August, or for approximately 4 months. Although the plan was changed to allow for more supergene/transition ore during the time of the machining work on the ball mill, it is not clear how this apparent change in mine plan and/or geological resource was realized.
	The mill performance for 2001 was subject to the disruptions from major repairs to one of the ball mills and other significant events that are not representative of future operations. Data produced since September 2001 are more representative of future mill operations now that repairs to the ball mill flanges have been completed on Mill A

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and replacement of flange bolts (larger bolts) in Mill B. The 2001 Kemess mill performance year to date is summarized in Table 16.1.

Table 16.1: Criteria for Mill Metallurgical Performance — January to October 2001

							Mill Head Grade						Recovery (%)						Concentrate
	Available		Crushed		Milled
Month	(%)		WMT		DMT		Cu (%)		Au (g/t)		Ag (g/t)		Cu		Au		Ag		DMT		Moist (%)		Cu (%)
January		82.5		1,244,222		1,219,338		0.280		0.926		1.448		80.62		67.74		58.28		12,140		9.18		22.64
February		82.6		1,115,751		1,093,436		0.279		0.900		1.286		80.74		67.77		68.38		11,730		8.85		20.98
March		82.9		1,254,787		1,229,691		0.308		0.973		1.362		85.94		71.02		70.19		14,722		8.43		22.11
April		82.4		1,162,004		1,138,764		0.239		0.752		0.978		79.33		64.50		58.35		10,049		8.07		21.45
May		90.2		1,491,979		1,462,139		0.170		0.962		0.943		63.00		62.91		56.58		5,893		8.79		26.63
June		73.1		1,135,391		1,038,654		0.342		0.975		1.102		77.30		67.08		63.10		11,274		8.39		24.35
July		78.9		1,532,704		1,305,932		0.173		0.899		0.875		73.23		63.78		53.26		6,307		7.87		26.27
August		84.2		1,454,423		1,458,574		0.249		0.832		0.991		76.91		64.10		58.88		11,991		7.64		23.32
September		70.3		1,052,679		1,097,452		0.240		0.686		0.894		78.51		62.73		54.43		9,960		7.62		20.80
October		84.9		1,510,971		1,478,398		0.281		0.914		1.113		78.92		65.26		63.46		15,011		8.14		21.85
November		90.5		1,458,721		1,430,841		0.262		0.743	N/A			76.70		64.70	N/A			13,528		7.83		20.17
December		—		—		—		—		—		—		—		—		—		—		—		—
Total		82.0		14,413,632		13,953,219		0.257		0.869		1.099		77.38		65.60		61.00		122,605		8.26		22.78

Note: November data included by AMEC from data Submitted by K. Stowe 12/07/2001

16.3	Metal Recovery
	Metal recovery for copper and gold are the major areas reviewed. Tables 16.2 and 16.3 outline the average operating performance expectations for milling operations estimated in March 2001 and actual performance as reported & major mill ore feed types.
	Table 16.2: Expected Mill Operations Metallurgical Performance Criteria — March 2001

	Recovery (%)
					Flotation		Concentrate
Ore Type	Copper		Gold		Grade (% Cu)		Moisture (%)
Hypogene		83		70		22		8
Supergene/Leach Cap >0.2% Cu		71		62		30-33		8-9
Supergene/Leach Cap <0.2% Cu		65		62		25-28		8-9
Transition		76		66		20-22		8-9

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Table 16.3: Actual Mill Metallurgical Performance Averages — 2001

	Recovery (%)
					Flotation		Concentrate
Ore Type	Copper		Gold		Grade (% Cu)		Moisture (%)
Hypogene		80.2		65.0		22.1		7.8
Supergene/Leach Cap 0.23% Cu		74.9		63.3		27.2		8.0
Supergene/Leach Cap 0.15% Cu		64.0		60.0		25.5		8.1
Supergene/Leach Cap Overall		67.3		61.0		26.5		8.1
Transition		77.2		64.5		22.2		7.9
Overall Weighted Performance		77.4		65.6		22.8		8.3

Note: Excludes zeros for days with no data reported in database because of downtime

The metallurgical performance of the supergene and transition ores has been as predicted and is well within the expectations set in March 2001. However, hypogene ores are expected to be the most predominant material mined over the course of operations, reportedly > 90%. A reconciliation of the 2001 operating data for hypogene versus the recovery projections in the Kemess life-of-mine (LOM) financial model (28 February 2001) is shown in Table 16.4.

Table 16.4: Comparison of Average Hypogene Mill Metallurgical Performance Predictions

Mine Life Avg.	Expected		2001 Mill Operating Data		Kemess LOM
Mill Ore Cu (%)		0.252		0.252		0.252
Mill Ore Au (g/t)		0.818		0.818		0.818
Cu Recovery (%)		84.5		80.2		85
Au Recovery (%)		71		65.0		73
Concentrate Cu (%)		24		22.7		22.9

Consistent and predictable metal recovery continues to be problematic for the hypogene material at Kemess. With slightly higher than average feed grades for both copper and gold, metal recovery for the hypogene ore was significantly lower than the second-half 2000 operating results and the performance levels predicted in March 2001. Applying AMEC’s March 2001 recommended performance curves for the mill operating data yielded recovery predictions that are generally 4% higher for copper recovery, 6% higher for gold recovery and 2% higher for copper concentrate grade. The Kemess LOM curves are more optimistic than the March 2001 curves. Plant availability problems are a significant detractor to making meaningful improvements in metal recovery. Hopefully these have now been resolved to a controlled and predictable level.

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	The higher sulphur/sulphide content in the hypogene, as reflected in higher pyrite (Py) content versus chalcopyrite (Cpy), will make it more difficult to control the flotation process and achieve a consistently optimal combination of recovery and concentrate grade (pyrite needing to be depressed, typically with lime, to the tails streams). It is not clear if this has contributed significantly to the lower recoveries. In addition, operations during circuit upsets due to modifications made to effect the major repairs programs and mill lining contributed to lower than expected performance.
	Several programs are in place at Kemess to find ways to improve operations. These have not been reviewed in detail with Kemess staff but are consistent with previous observations and discussions. These programs should produce some benefits over time and include flash flotation and Knelson concentrators for gold and native copper recovery; speeding up the rougher mechanisms; additional modifications to the cleaner circuit, including column flotation, thickener improvements and coagulant changes; and improving the sampling equipment and methodology.
	At this time, recovery expectations for supergene and transition ores are the same as those given in the March 2001 report. Without sustained proven performance to the contrary, AMEC believes metal recoveries may be more conservative, especially given recent performance. Anticipated copper recovery from hypogene may not be higher than 80% to 82% at the average copper ore grade of the deposit. Anticipated gold recovery may be lower at 65% to 66%.
16.4	Plant Availability
	Operating data for 2001 shows an overall availability of approximately 82%. These results are less than the life of mine plan of 90% and are less than the average monthly availability of 88.4% (annualized basis) that was expected by AMEC in March 2001. The lower than expected availability results are due in large part to higher than planned SAG and Ball mill relining frequency during the period, a few one-time modifications events in the tailings system; main powerline supply; Ball Mill A flange face machining and flange bolt replacement of Ball Mill B, and continuing problems with ball mill and SAG liners/bolts over the course of the year. Most of these are similar in nature or the same as was previously noted in March 2001. Now that the ball mill flanges have been addressed, an overall increase in average availability of 3% (as reported in March 2001) to 88% should be achievable. The average availability achieved in 2001 during the October/November period was 87.7% with November at 90.5%. Winter operating problems and their impact on plant availability have been reduced through an aggressive winterization plan implemented in 2000; as such, this is no longer a significant area of concern. While there were some reports of downtime

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	due to frozen ore jamming the crushing circuit and other winter-related issues, no significant downtime is attributable solely to winter operations.
	Based on the current data, the March 2001 availability expectation is optimistic and should be revised to an average of 88% to account for the maintenance problems that are likely to continue as the equipment ages. Month-to-month availability will vary in accordance with the major mill relining work and other planned maintenance activities. Given that all the mills were relined in the period, this too contributing to lower average, an average of 88% availability is still reasonably achievable if the October/November 2001 data is now more indicative of normal availability.
16.5	Mill Throughput
	Kemess reported an average mill throughput of 41,776 t/d at 82% overall availability in 2001. Had the predicted availability (88.4%) been achieved, the milling rate would have been approximately 45,000 t/d, which is lower than the Kemess life-of-mine plan level of 48,000 t/d. Other than availability, the shortfall in milling throughput is partially explained by the mill reconfiguration and lower tonnages that were processed to allow for flange machining on Ball Mill A and Mill B bolt replacement, but not entirely. Apparent operation at less than design maximum milling rates for the respective feed ores during mill running time has not been fully investigated at this time. November 2001 milling rates approach the expected level, (reported average of 47,678 t/d). Hypogene throughput averaged 41, 137 t/d at 80.4% availability. Had availability been the predicted 88.4%, the milling rate would have been approximately 45,230 t/d. Higher than expected ore hardness was attributed by Kemess for some of the lower milling rates. Estimates for hypogene ore based on 2001 data suggest a throughput of 46,060 t/d as a maximum rate for 90% availability. Supergene/leach cap throughput averaged 41,395 t/d at 80.3% availability. Had availability been at the predicted 88.4%, the milling rate would have been approximately 45,570 t/d. This is considerably lower than the rate of 49,250 t/d anticipated in March 2001.
	Transition ore throughput averaged 42,484 t/d at 86.8% availability. Had availability been the predicted 88.4%, the milling rate would have been approximately 43,270 t/d. Milling rates improved significantly at the end of November 2001 were averaging approximately 50,000 t/d and are more indicative of previous performance and probable future expectations.
	The 2001 data therefore show that Kemess may not achieve the life-of-mine plan throughput rate of 48,000 t/d, even at the plan target of 90% availability, by operating at the throughput rates achieved in 2001 with mainly hypogene ore. Based on the current operating conditions, a milling rate of 44,500 t/d is achievable at the

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	recommended 88% availability under the assumption that hypogene ore is the predominant material remaining to be mined (estimated to be >90%). This is less than the anticipated milling rate of 47,650 t/d for hypogene in March 2001. These figures will increase proportionately if significant amounts of transition and/or supergene materials are yet to be mined.
	Northgate’s management still believes there are opportunities to improve throughput and that the life-of-mine plan throughput rate of 48,000 t/d can be achieved.
16.6	Mill Flotation Concentrate
	Concentrate data for 2001 continued to show very consistent moisture results, averaging 8.3% with no excursions 10%. This represents a considerable achievement by Kemess that is expected to continue, as it is apparent that past issues concerning concentrate moisture levels have been resolved. Concentrate quality remains in the acceptable range, at 26.5% Cu for supergene ore campaigns and 22.1% Cu for hypogene ore campaigns. It should be noted that a significant amount of transitional ore was also processed during the supergene campaigns.
16.7	Conclusions
	The usefulness of the historical performance data from the mill operations in 1999 and early 2000 was compromised by poor metallurgical control, sub-par operating practices and extremely low equipment availabilities due to mechanical failures. Reliability, performance and overall control have improved significantly since June 2000, and the data collected subsequently are considered to be sound, contiguous and substantial enough to make responsible predictions about future milling operations.
	The current performance data demonstrate the benefits realized from design improvements and better attention to the metallurgical operating performance of the mill. However, there is too small an amount of data (2 months) since the final significant mill improvements and disruptions (mill re-linings) were completed and this is not sufficient to establish a clear and definitive performance basis for the longer term future. The October/November 2001 data does indicate the current status and potential for improved future results.
	Revised estimates for evaluating future Kemess milling performance are summarized in Table 16.5.

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Table 16.5: AMEC Estimated Criteria for Mill Metallurgical Performance — November 2001

	Recovery (%)
					Flotation		Concentrate
Ore Type	Copper		Gold		Grade (% Cu)		Moisture (%)
Hypogene		80-82		65-66		22		8
Supergene/Leach Cap > 0.2% Cu		75		62		28		8-9
Supergene/Leach Cap < 0.2% Cu		64		60		25		8-9
Transition		76		65		22		8-9

The estimated average mill throughput is 44,500 t/d hypogene ore at an 88% average plant availability. Supergene ore throughput should average 50,000 t/d if mill feed rates achieved in October/November are sustained.

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17.0	MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES
	The mineral resource and reserve estimates for the Kemess South and Kemess North deposits were calculated under the direction of Mr. Michael Hibbitts, Northgate’s Qualified Person for mineral resources and reserves. The estimates were made from 3-dimensional block models utilizing commercial mine planning software (MineSight®). As part of MRDl’s independent review, grade interpolation parameters were examined for suitability and the models checked for validity. The resource and reserve calculation method and appropriateness of resource and reserve classification categories used were also examined.
17.1	Geologic Models
	The Kemess deposits exist in two geologic block models, Kemess South and Kemess North. Limits and cell block sizes for each MineSight® project are shown in Table 17.1.
	Table 17.1: Project Limits and Model Cell Block Sizes

		Minimum		Maximum		Block Size
Deposit	Axis	(m)		(m)		(m)
	X (east)		8,800		11,000		20
Kerness South	Y north)		9,100		10,600		20
	Z (elev.)		600		1,440		15
	X (east)		8,000		12,000		20
Ketness North	X (east)		15,000		17,000		20
	Z (elev.)		1,000		2,005		15

Modelling philosophy was similar for both models, where geologic features were important in controlling the extent of interpolation. The geologic features of the Kemess South deposit are well known, and eight lithologic units are recognized. These are named differently depending on where they are located relative to the north-south fault in the pit. For example, unit 3 becomes unit 13 to the east of the fault. Units used in grade interpolation are listed in Tables 17.2 and 17.3.

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Table 17-2: Geology Unit Names and Block Model Codes — Kemess South

Code	Mineralized Geology Unit
1 2 3 4 5 6 7 8	Overburden Leach Cap Monzonite Supergene Monzonite Barren Dyke Hypogene Mineralization Takla Volcanics Takla Sediments Tertiary Sediments

Table 17.3: Geology Unit Names and Block Model Codes — Kemess North

Code	Mineralized Geology Unit
1 2 3 4 5 6 7 8 9	Overburden Post-Ore Mafic Dyke Mafic flow Post-Ore Feldspar Porphyritic Syenite Dykes Hypogene Mineralization Takla Volcanics Bladed Feldspar Porphyry Monzodioritic Dykes Faults, Breccias, Shears, Gangue

The sectional geological outlines for the Kemess South deposit were used for tagging geology codes in the assay database. MRDI checked the results of the geology tagging and found it to have been done properly.

The assays were composited into 15 m benches, honouring geology units. MRDI reviewed the results of the compositing process and found it to have been done correctly. However, because of the way MineSight® electronically stores this type of data, the resulting composite database is sensitive to selection protocols. As long as the composites are selected by geology codes, the proper composites are chosen in grade interpolation.

The Kemess North resource model was built with 5 m composites and a bench height of 15 m.

Mine staff conducted spatial analyses of composites from Kemess South drill hole data using multi-directional variograms. MRDI reviewed this work and found that the interpreted models were reasonable. Variograms for Kemess North were recalculated

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by MRDI and found to define the same trend as introduced in the grade modelling by Northgate. Variogram parameters for copper and gold are shown in Table 17.3 for the hypogene mineralization at Kemess Mine. It should be noted that the dip and plunge interpreted from the variograms vary from the east side (rock type 5) of the north-south fault to the west side (rock type 15). East of the fault the plunge and dip are respectively 12° and -20. West of the fault the plunge is -10° and the dip -8°.

Table 17.3: Variogram Parameters — Kemess South

Geology					Sill		Range Y		Range X		Range Z		Rotation Z
Codes		Type	Nugget		Difference		(ft)		(ft)		(ft)		Left Hand
Copper
	5	Sph		0.002		0.008		287		887		40		90
	15	Sphl		0.003		0.002		52		42		20		90
		Sph2				0.009		387		181		100		90
Gold
	5	Sphl		0.012		0.002		120		179		30		90
		Sph2				0.007		234		405		87		90
	15	Sphl		0.010		0.006		47		80		10		90
		Sph2				0.010		166		292		64		90

Note: Sph = spherical model

For Kemess North the hypogene mineralization was assigned grade using the variograms listed in Table 17.4. Separate variograms were used from the top of the block model down to 1,600 m elevation, as the orientation of the hypogene mineralization varies below this level.

Table 17.4: Variogram Parameters — Kemess North

Geology					Sill		Range Y		Range X		Range Z		Rotation Z
Codes		Type	Nugget		Difference		(ft)		(ft)		(ft)		Left Hand
Copper
	5	Sph1		0.002		0.006		226		295		37		70
		Sph2				0.009		276		477		70		70
5 (below 1600 m)		Sph1		0.001		0.005		126		192		38		50
		Sph2				0.007		214		585		116		50
Gold
	5	Sph1		0.007		0.018		137		216		31		80
		Sph2				0.023		280		514		121		90
5 (below 1600 m)		Sph		0.008		0.028		201		424		150		50

Note: Sph = spherical model

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The specific gravity database used for the current estimate at Kemess South includes 3,162 measurements done on pulps from the El Condor 1991 drilling program. Subsequently, selected representative core samples were sent to Intertek Testing Services (Bondar-Clegg) in Vancouver for SG determination using the wax immersion method. The selected core included 229 samples from the supergene zone, 45 samples from the hypogene and 2 samples from the Sustut Group rocks that overlie the west half of the deposit. Kemess staff used average specific gravities as per Table 17.5.

Table 17.5: SG Values for Selected Core Samples

	Used in		Average from		Number of		Difference
Zone	Block Model		Testing		Samples		(%)
Supergene monzonite		2.69		2.58		229		4.1
Hypogene monzonite		2.71		2.62		45		3.3
Sustut rocks		2.56		2.39		2		6.6

	The SG measurements on core indicate that the supergene and hypogene ore tonnages, and therefore reserves, may be overestimated by 4% and 3%, respectively. The small number of determinations in the Sustut rocks potentially indicates that waste tonnages may also be overestimated. It should be noted that these percentages fall generally within the margin of error for the reserve calculation and that MRDI does not recommend that the tonnages shown in the life-of-mine plan be changed. An average specific gravity of 2.62 was used for the Kemess North resource model.
17.2	Interpolation Plans
	Values for copper and gold ware interpolated by ordinary kriging. At Kemess South all domains were treated as hard boundaries, which means that only composite values of the same ore type or geology code as the model block could be used to interpolate that block’s grades. For the Kemess North model only one large mineralized zone was interpreted, as only hypogene mineralization has been identified to date.
	Kemess South Plan o    Minimum composites to interpolate a model block = 3 o    Maximum composites to interpolate a model block =16 o    Maximum X (east) search = 390 m o    Maximum Y (north) search = 180 m o    Maximum Z (elevation) search = 100 m.

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	Kemess North Plan o    Minimum composites to interpolate a model block = 3 o    Maximum composites to interpolate a model block = 16 o    Maximum X (east) search = 477 m o    Maximum Y (north) search = 276 m o    Maximum Z (elevation) search = 70 m.
17.3	Model Validation
	MRDI completed a reasonableness review of the Kemess block models, looking primarily at the grade interpolation relative to drill hole composite values. This was done mainly by inspecting sections and plans. Overall good agreement exists between the composite and model block copper and gold values. Examples of representative Kemess South and North sections and plans, containing block model copper and gold grades, drill hole composite values, geology unit outlines and ultimate pit outline, are included in Appendix B.
	To check the smoothing in the block model, MRDI constructed change-of-support models. To ensure that the change-of-support models were appropriate for validating the current set of model estimates, MRDI updated nearest-neighbour estimates and variogram parameters. MRDI calculated block variance factors of 0.76 for copper and 0.92 for gold for the change-of-support comparison, based on Kemess’ 20 m x 20 m mining block size. These are high block variance factors, indicating good spatial continuity for this type of deposit.
	The distributions of calculated 20 m x 20 m x 15 m blocks for copper and gold are shown by dashed lines on the grade-tonnage curves in Figures 17.4 and 17.5. These nearest-neighbour results represent the grade distributions, based on this block size, obtained from the change-of-support models. The solid lines in the figures show the grade-tonnage distributions obtained from the block estimates. The grade-tonnage predictions produced for the Kemess models show that grades and tonnages are validated by the change-of-support calculations.
	The change-of-support analysis confirms that the copper and gold models adequately incorporate allowances for internal dilution in the range of metal cutoff grades used at Kemess. External dilution and ore losses are considered in the mine planning section of this report.
	Notes for the following graphs: Herco Model is a polynomial model                                                       OK Model is the ordinary kriged model

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Figure 17.1: Grade-Tonnage Curves for Copper

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Figure 17.2: Grade-Tonnage Curves for Gold

For Kemess North MRDI also checked the block model estimates for global bias by comparing the average values (with no cutoff) from the model (ordinary kriging) with the means from nearest-neighbour estimates (the nearest-neighbour estimator produces a theoretically unbiased estimate of the average value when no cutoff grade is imposed and is a good basis for checking the performance of different estimation methods).

To conduct this check, nearest-neighbour copper grades were estimated and compared to the existing model kriged copper grade. Mean and standard deviation values from ordinary kriging and nearest-neighbour copper estimates (at zero copper cutoff grade) are tabulated in Table 17.6. Results show no evidence of bias in the estimates.

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Table 17.6: Copper Mean and Standard Deviation (s.d.)
Values from Ordinary Kriging and Nearest-Neighbour Estimates

Kemess North
Hypogene	Type	Mean%		s.d
Copper	Krige		0.1023		0.0789
	NN		0.1035		0.1085
Gold	Krige		0.2115		0.1311
	NN		0.2104		0.1938

17.4	Resource Classification and Summaries
	The mineralization at the Kemess Mine as of 31 December 2001 is classified into Mineral Reserves and Resources and is shown in Table 17.7.
	Table 17.7: 2001 Kemess Mineral Reserves and Resources

					Grades
	Cutoff Grade								Contained Metals
	Gold (gmt)				Gold		Copper		Gold		Copper
	Equivalent		Tonnes		(gmt)		(%)		(oz)		(M lb)
Inferred Mineral Resource
		0.6		442,000,000		0.40		0.23		5,700,000		2,200
Kemess North		0.8		170,000,000		0.50		0.29		2,700,000		1,100
Mineral Reserve Kemess South				145,911,000		0.65		0.24		3,063,000		756

Based on the geostatistical work performed by MRDI, the Kemess South deposit would require drilling on 70 m centres to classify the resource as measured. The indicated category would encompass the material defined by a 100 m x 100 m drill pattern. These grids are consistent with the patterns used for most large porphyry copper projects such as Quebrada Blanca, Escondida Norte and Cerro Colorado. Typically, large porphyries are drilled on 50 m to 70 m centres for measured/proven and on 75 m to 200 m centres for indicated/probable, depending on the grade variability.

The entire Kemess deposit within its ultimate pit outline is classified as Probable Mineral Reserve. This level of reserve category is supported by good mill-blast hole-model reconciliation and an adequate drill pattern (100 m x 100 m). Economic parameters are also well established and supported by current production.

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All of the Kemess North deposit is classified as an Inferred Resource to reflect the lower drill hole density and resultant higher uncertainty in grade and geologic continuity. Limited metallurgical testwork has been undertaken for Kemess North (see Section 16) but no economic evaluation has been completed as yet.

The mine site personnel and Knight Piesold base the slope designs for the Kemess South open pit on ongoing geotechnical studies. The current design has been broken into nine design sectors with inter-ramp angles ranging from a low of 35° to a high of 47°. The mine staff collect, process and interpret the geotechnical monitoring data. Reports on the status of the pit slope are compiled regularly and discussed with operations personnel. An independent geotechnical engineer also reviews the data on a quarterly to semi-annual basis.

The cutoff NSR value of the Kemess South deposit reserves is Cdn $4.38 for pit elevations above 1,290 m. Below 1,290 m elevation, the cutoff NSR value increases by increments of Cdn $0.01 per level down to the bottom of the pit at 1,005 m. This calculation is based on Northgate’s expectations for:

o Copper prices

o Copper recovery

o Copper smelting, refining and shipping costs

o Gold prices

o Gold recovery

o Silver prices

o Royalties

o U.S. to Canadian dollar exchange rate.

The mine staff review the copper NSR calculation on an ongoing basis and update the equation as required.

MRDI has checked the NSR calculation and determined that a profit would be made at the expectations summarized in Table 17.8.

Table 17.8: Price, Recovery and Concentrate Grade — Kemess South

Commodity	Price (U.S.$)	Recovery (%)		Concentrate Grade (%)
Copper	0.95/lb		82		23
Gold	300 oz		70

Metal prices are based on Northgate’s long-term price projections, and are reasonable and within market parameters (see Section 19).

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Additional parameters considered in calculating the NSR value are based on actual concentrate distributions and costs; these are trucking, shipping and payments. Starting in February 2002, all the concentrate produced by the Kemess Mine will be shipped to Noranda’s Horne smelter in Rouyn-Noranda, Quebec.

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18.0	OTHER DATA AND INFORMATION
18.1	Reconciliation
	Reconciliation of block model estimates to actual production at operating mines is an important measure of the predictive nature of the block model for future production. Two sets of data are available from the Kemess South deposit to compare with the block model estimates: blast hole data and milled production data. MRDI evaluated these data to obtain a level of confidence in block model performance.
	MRDI reviewed monthly production data for September 1998 and reconciliation data for the year 2000, all provided by Kemess. Year 2001 reconciliation data were also reviewed but treated separately.
	Reconciliation is the assessment of the ability to estimate in-situ resources and the ability to mine those resources. For a reconciliation assessment, MRDI requires that the following factors be calculated:

o Assessment of the ability to estimate in-situ resources (F1): A check on the accuracy of the Kemess block model in predicting annual quantities of in-situ mineable reserves (tonnage, grade, pounds of copper and ounces of gold) for both mill and stockpile material. The accuracy is summarized by the F1 factors, which are calculated for tonnes and grades:

F1 = Actual In-Situ Quantities — Blast Hole Model
Estimated In Situ Quantities — Diamond Drill Hole Model

o Assessment of the ability to mine in-situ reserves (F2): An evaluation of the effects of misclassification (dilution and/or ore loss) on the in-situ mineable reserves. This effect is summarized by the F2 factors, which are calculated for tonnes and grades:

F2 = Run-of-Mine Quantities - Mill Data
Actual In-Situ Quantities - Crusher Data

F1 and F2 factors can be multiplied together to yield F3 factors, which assess the impact of the combined effect of estimation errors and selection errors to predict recoverable reserves:

F1 x F2 = F3 = Run-of-Mine Quantities - Mill Data
Estimated In-Situ Quantities

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MRDI calculated the F1, F2 and F3 factors for the Kemess South deposit from the monthly data provided by Kemess. The factors by year are summarized in Table 18.1.

Table 18.1: Summary of F1, F2 and F3 Factors for Kemess South, 1998 to 2000

		Tonnes		Cu (%)		Au (g/t)
F1	1998		1.064		0.992		1.027
	1999	1.000			0.962		0.991
	2000	1.000			1.041		0.995
	Overall		1.012		1.000		0.999
F2	1998		0.764		0.905		0.990
	1999		1.014		0.935		1.017
	2000		0.986		0.919		1.049
	Overall		0.959		0.923		1.025
F3	1998		0.808		0.895		1.014
	1999		1.014		0.898		1.004
	2000		0.990		0.946		1.041
	Overall		0.966		0.917		1.021

Note: The low F2 factor for 1998 reflects low milling tonnage in December 1998

Only the year-end 2000 reconciliation is considered to be complete and accurate, because it is based on the final smelter return data, without the crusher feed grades generated from daily advances.

The final year-end 2000 reconciliation, based on the smelter returns, is shown in Table 18.2. This overall reconciliation provides a glimpse at a steady-state operation and indicates a shortfall of 2% in copper between the drill hole model and the milled material. The difference between the blast hole model and the milled ore is 6% for copper and 1% for gold. These differences are consistent with variation patterns seen from month to month and are normal fluctuations for a large porphyry gold-copper mine such as Kemess.

Further review by MRDI of year 2001 production data available to date reveals a different pattern in the milled gold grades. The block model monthly estimates were calculated by mine staff using month-end topographic surfaces. The milled material figures contain the month-end tonnage and grade. The ratio of milled material to block model is then a measure of how well the milled ore tonnage and grade reconcile to predictions made by the block model. The results, shown in Table 18.3, demonstrate that the block model does an adequate job of predicting copper grade but underestimates the gold grade by an average of 15% on a monthly basis.

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Table 18.2: Year-End 2000 Reconciliation (provided by Kemess)

	Tonnes			Cu (%)		Au (g/t)
Blast Hole Model vs. Diamond Drill Hole Model
2000 BHS Model (Final Reconciliation)		14,700,023			0.234		0.785
2000 DDH Model (Final Reconciliation)		14,700,023			0.225		0.786
Difference (BHS vs. DDH)		0			0.009		-0.001
Difference in Percentage (BHS vs. DDH)		0.00	%		0.039		-0.001
Mined vs. Milled
2000 BHS Model (Final Reconciliation)		14,700,023			0.234		0.785
2000 DDH Model (Final Reconciliation)		14,700,023			0.225		0.786
2000 Milled (Final Reconciliation)		14,426,982			0.220		0.778
Net gain in stockpile		310,416			0.227		0.794
Total Milled and Stockpiled		14,737,398			0.220		0.778
Difference (DDH Model vs. Milled)		37,375			-0.005		-0.008
Difference in Percentage (DDH Model vs. Milled)		0.25	%		-2.16		-0.98
Difference (BHS Model vs. Milled)		37,375			-0.01		-0.007
Difference in Percentage (BHS Model vs. Milled)		0.25	%		-5.98		-0.85

Table 18.3: Year 2001 Copper Reconciliation Data

									Mill vs. DDH		Mill vs. BHS
	Tonnes								Difference		Difference
	Milled		DDH		BHS		Milled		(%)		(%)
January		1,219,338		0.280		0.282		0.280		0.0		-0.7
February		1,093,436		0.262		0.266		0.279		6.5		4.9
March		1,229,691		0.251		0.276		0.308		22.7		11.6
April		1,138,764		0.230		0.221		0.239		3.9		8.1
May		1,462,139		0.182		0.174		0.170		-6.6		-2.3
June		1,038,654		0.316		0.326		0.342		8.2		4.9
July		1,305,932		0.198		0.274		0.173		-12.6		-36.9
August		1,458,574		0.257		0.239		0.249		-3.1		4.2
September		1,097,452		0.200		0.251		0.240		20.0		-4.4
October		1,478,398		0.279		0.279		0.279		0.0		0.0
Total		12,522,378		0.244		0.257		0.253		3.5		-1.5

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Table 18.4: Year 2001 Gold Reconciliation Data

									Mill vs. DDH		Mill vs. BHS
	Tonnes								Difference		Difference
	Milled		DDH		BHS		Milled		(%)		(%)
January		1,219,338		0.834		0.921		0.926		11.0		0.5
February		1,093,436		0.737		0.884		0.900		22.1		1.8
March		1,229,691		0.743		0.950		0.973		31.0		2.4
April		1,138,764		0.726		0.781		0.752		3.6		-3.7
May		1,462,139		0.906		0.962		0.962		6.2		0.0
June		1,038,654		0.911		0.910		0.975		7.0		7.1
July		1,305,932		0.818		0.924		0.899		9.9		-2.7
August		1,458,574		0.730		0.790		0.832		14.0		5.3
September		1,097,452		0.522		0.681		0.686		31.4		0.7
October		1,478,398		0.733		0.828		0.899		23.2		8.6
Total		12,522,378		0.768		0.865		0.883		14.9		2.1

The benches mined in year 2001 show systematically higher gold grades than those predicted by the drill hole model. It is MRDI’s opinion that a number of factors may explain the difference. Further work is required to determine if the same trend will continue and the extent of the potential grade increase.

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19.0	REQUIREMENTS FOR TECHNICAL REPORTS ON PRODUCTION
19.1	Mining Operations
	The Kemess Mine currently operates from one open pit. The pit is mined on 15 m benches and is split into three phases. The final wall is designed with a combined single- and double-bench configuration. The wall angles are noted in Section 17 of this report and documented in Appendix E. The pit ramp widths are 35 m and exceed the design width as defined by the Health, Safety and Reclamation Code for Mines in British Columbia for the equipment utilized at the operation. All roads have been designed for a maximum grade of 10%.
	Major mining equipment includes the following:

o
o
o
o
o
o Euclid R260 haulage trucks
1 P&H 2800XPB rope shovel
Hitachi EX3500 hydraulic front shovel
1 LeTourneau L-1400 front-end loader
1 Gardner Denver GDCM blasthole drill
1 P&H XP 100 blasthole drill.

The ore is hauled directly to a gyratory crusher and coarse ore stockpile area south of the pit. The crushed ore is fed to two semi-autogenous (SAG) grinding mills, followed by two ball mills and one regrind mill. Flotation circuits are then used to produce a gold-copper-silver concentrate; a process mill flow sheet is included in Appendix E.

Mill tailings are pumped through two 7,500 m long, 66 cm diameter lines to the tailings facilities. Tailings dam construction is an ongoing project over the life of the mine, with monitoring and design work being performed by AMEC E&E from Burnaby, B.C.

The concentrate is trucked in bulk approximately 380 km via gravel road to a rail spur at Mackenzie, B.C. The concentrates are then loaded onto gondola railcars and transported to the Port of Vancouver, where they are transferred to ocean-going vessels for shipment to Japan.

The mine staff have prepared a long-term production forecast that is designed to minimize operating costs while maximizing mill throughput. The staff then calculated haulage productivities based on assumed profiles for the production forecast. This is combined with the associated shovel productivities to form the basis of the long-range plan cost forecast.

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The cutoff grade for the Kemess Mine is based upon the projected operating costs and royalties payments for the mine. This is then compared to the calculated Net Smelter return (NSR) for the ore material. Based on this forecast, the Kemess pit reserve will be depleted in 2009 (see Table 19.1).

Table 19.1: Production Forecast

	Tonnes (000) of Ore Mined
Area	2001		2002		2003		2004		2005		2006		2007		2008		2009
Kemess South		17,173		17,500		17,906		17,520		17,867		17,827		17,787		17,528		3,588
Total Mined		36,711		47,513		47,897		47,725		49,242		41,869		29,194		19,038		3,745

	The ore defined for this production forecast includes only that material that exceeds the NSR cutoff.
19.2	Process Metal Recoveries
	This topic has been addressed in detail in Section 16.
	Metal recoveries for evaluating future Kemess milling performance are summarized in Table 19.2.
	Table 19.2: AMEC Estimated Criteria for Mill Metallurgical Performance - November 2001

	Recovery (%)
					Flotation		Concentrate
Ore Type	Copper		Gold		Grade (% Cu)		Moisture (%)
Hypogene		80-82		65-66		22		8
Supergene/Leach Cap >0.2% Cu		75		62		28		8-9
Supergene/Leach Cap <0.2% Cu		64		60		25		8-9
Transition		76		65		22		8-9

19.3	Markets
	Northgate markets the Kemess concentrate, all of which is currently sold to Glencore Limited, for shipment to Japan. Starting in February 2002, all concentrates will be sold to Noranda Inc.’s Horne Smelter in Quebec.

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19.4	Contracts
	As noted above, all concentrates produced until February 2002 will be sold to Glencore and shipped to various smelters in Japan and elsewhere. The smelter contract contains provisions for penalty payments for deleterious elements, but to this point none has been exercised.
	The new contract to the Horne Smelter is for a two-year period, with an option for a third year. The contract includes provisions for re-negotiating the shipping costs if a road is built between the mine site and the Port of Stewart, B.C.
19.5	Environmental Considerations
	The current reclamation bond for the Kemess Mine has been set at Cdn $12 million and is in good standing. Reclamation and remediation are ongoing at the mine site and are budgeted items.
	The mine staff produce an annual reclamation report, as required under B.C. law. This report documents areas of disturbance, future development and reclaimed ground. The mine staff also prepare an annual spreadsheet documenting liabilities at the end of each year, which is submitted to the government. The latest filed reclamation report was for year 2000.
19.6	Taxes
	The operation is currently subject to the following tax: o British Columbia mineral tax (3% on resource income). The above tax has been included with the economic model.
19.7	Capital and Operating Cost Estimates
	Mine staff developed the 2001 budget operating costs from first principles. Annual equipment utilization hours were derived from the forecast, then costed based on current unit operating costs. Minor support functions were estimated based on unit per mined tonne values and added to the overall costs to produce a mine operating cost. The year-over-year costs were incremented to reflect the increasing depth of the pit. The mill and the site general and administrative costs were budgeted based upon the material movement for 2001, then assumed to be fixed on a dollar per tonne basis for the life of mine. Table 19.3 summarizes the mine operating costs for the duration of the existing mine life.

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Table 19.3: Projected Operating Costs

Function	Operating Cost (Cdn$/t ore milled)
Mining		2.01
Milling		2.47
General and Administrative		1.07
Total		5.64

	MRDI also reviewed the projected capital expenditures. Capital expenditures include sustaining capital for raising the ailing dam, plant modifications, camp expenditures and incremental mining equipment. Total sustaining capital over the life of the mine has been estimated at $112 million, of which the tailing facility comprises 65%.
	The operating costs show a general increase throughout the life of mine; MRDI believes this reflects the anticipated conditions. The capital expenditures reflect the increased haulage distances (purchase of five 240 t trucks over five years) and work required to raise the tailings facility to accommodate the mill tailings.
19.8	Economic Analysis
	Northgate has completed an after-tax economic analysis for the property using the long-term production forecast as a base. The assumptions incorporated within the model are as follows:
	Interest Rate

o 8.5%

Exchange Rate

o Cdn $1.45 to U.S. $1.00

Metal Prices

o Copper                  U.S. $0.95/lb

o Silver                     U.S. $5.00/oz

o Gold                      U.S. $300/oz.

The above assumptions are higher than current market prices for the three metals but lower than the 10-year average prices. Interest rates and the exchange rate are currently more favourable than these assumptions. Over the past several years most of the major gold producers have restated their reserves utilizing a lower gold price.

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After-tax cash flows for the mine calculated by Northgate are positive, confirming that reserves can be expected to generate a profit with the assumptions used in the model.

Sensitivities to the model have been run on variations of the following items:

o Copper price

o Gold price

o Exchange rate.

	Northgate regards the details of the economic and sensitivity analysis as confidential. The after-tax economic analysis has been reviewed, and MRDI believes it to be sound and to normal industry standards.
19.9	Payback
	Currently the mine has received a U.S. $100 million loan from the Bank of Nova Scotia. At the above stated parameters, the loan is to be paid down by September 2007.
19.10	Mine Life
	Proven and probable mineral reserves are projected to sustain the mining operation until 2009.

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20.0	INTERPRETATION AND CONCLUSIONS
	MRDI reviewed pertinent geological, mining and metallurgical data from the Kemess Mine to obtain a sufficient level of understanding to assess the Kemess 2000 Mineral Reserve and Resource statement. Following is a list of MRDl’s general conclusions:

o The geology of the Kemess South porphyry gold-copper deposit is well understood. The type and extent of the mineralization are well defined. The Company has appropriately integrated that knowledge into the geologic block models, mining practice and metallurgy.

o The database to support the Mineral Reserve and Mineral Resource comprises 334 drill holes. The high quality of the assay data used to estimate the Mineral Reserve at Kemess (South deposit) is supported by good reconciliation of material milled to block model grades and tonnages. MRDI verified the database used for the block model estimates for the Mineral Reserve by a 5% check to original assay and survey data.

o The 2000 block model was developed using industry-accepted methods. MRDI validated the model estimates and found them to reasonably estimate grade and tonnage for the Kemess South and Kemess North deposits. Based on actual production figures for year 2001, the gold grade estimated for at least some parts of Kemess South may have been understated by approximately 15%.

o The Mineral Reserve and Mineral Resource at Kemess is classified into Probable Reserve (Kemess South) and Inferred Resource (Kemess North). MRDI assessed the criteria used by Northgate for this classification and agrees with them.

o The cutoff grade strategy employed by Kemess is sound. MRDI reviewed the various assumptions that made up the basis for the cutoff value and found them to be sound and based on industry-accepted parameters.

o MRDI found that Kemess metallurgical expectations are reasonable and based on stable metallurgical results during recent production. MRDI found the operating cost estimate to be reasonable and to have been calculated using sound industry-accepted practices.

o MRDI found the assumptions used for the economic forecast to be within market parameters and to be valid assumptions for an economic forecast.

This independent review by MRDI supports the 2000 Kemess Operation Mineral Reserve and Mineral Resource statement.

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21.0                      REFERENCES

     Canadian Mineral Deposit Types: A Geological Synopsis, O.R. Eckstrand, Geological Survey of Canada, Economic Geology Report 36, 1984.

     Knight Piesold Ltd. (1993a). “Report on Tailings Facility Surficial Geotechnical Investigations”, March 1.

     Knight Piesold Ltd. (1993b). “Prefeasibility Study, Tailings Storage Facility”, June 1.

     Knight Piesold Ltd. (1994a). “Report on Project Water Management”, January 1.

     Knight Piesold Ltd. (1994b). “Report on Tailings Storage Facility Surface and Groundwater Management”, January 1.

     Knight Piesold Ltd. (1994c). “Report on Tailings Storage Facility Addenda to Mine Development Certificate Applications”, May 26.

     Knight Piesold Ltd. (1995a). “Overview Report and Responses to Agency Comments”, March 1.

     Knight Piesold Ltd. (1995b). “Report on Tailings Storage Facility Supplementary Geotechnical Investigations and Updated Design Basis”, March 1.

     Knight Piesold Ltd. (1996a). “Tailings Storage Facility Report on 1996 Geotechnical Investigations”, June 25.

     Knight Piesold Ltd. (1996b). “Tailings Storage Facility Interim Design Report”, September 18.

     Knight Piesold Ltd. (1996c). “Tailings Storage Facility Preliminary Information package for Landslide on Right Abutment”, December 5.

     Knight Piesold Ltd. (1996d). “Supplementary Geotechnical Information Summer/Fall 1996 Site Investigation”, December 20.

     Knight Piesold Ltd. (1997a). Tailings Storage Facility Supplementary Design Report”, February 28.

     Knight Piesold Ltd. (1997b). “Tailings Storage Facility Final Design Report”, August 18.

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     Knight Piesold Ltd. (1997c). “Information Summary for I.R.P.”, September 15.

     Knight Piesold Ltd. (1998a). “Tailings Storage Facility Summary of Tailings Embankment Instrumentation”, April 8.

     Knight Piesold Ltd. (1998b). “Technical Specifications for Stage 2 Tailings Embankment Raise”, April 30.

     Knight Piesold Ltd. (1998c). “Tailings Embankment Stability Analyses”, May 5.

     Knight Piesold Ltd. (1998d). “Tailings Storage Facility, Report on Stage 1 Construction”, November 30.

     AGRA Earth & Environmental Limited (1998a). “Kemess South Tailings Impoundment — Review of Design Flood Event and Freeboard”, December 17.

     AGRA Earth & Environmental Limited (1998b). “Kemess Tailings Dam Instrumentation Review”, December 18.

     AGRA Earth & Environmental Limited (1999a). “Kemess Tailings Dam: Design of Stage 3 Raise”, April 14.

     AGRA Earth & Environmental Limited (1999b). “Kemess Mine Tailings Storage Facility: Stage 2 As-Built Report and Annual Review”, June 14.

     AGRA Earth & Environmental Limited (1999c). “Instrumentation and Laboratory Testing Programs: Kemess Tailings Dam”, December 23.

     AGRA Earth & Environmental Limited (2000a). “Kemess Mine Tailings Storage Facility — Design of Stage 4 Raise of tailings Dam”, May.

     AMEC Earth & Environmental Limited (2000b). “Kemess Mine Tailings Storage Facility — Stage 4 Construction As-Built Report”.

     AMEC Earth & Environmental Limited (2000c). “Kemess Mine Tailings Storage Facility — Stage 5 Tailings Dam Raise, Design Report”, December 18.

     GRP (1996). “Report No. 1 of Geotechnical Review Panel, Tailings Impoundment Design, Kemess South Project”, April 12.

     GRP (1997a). “Tailings Program — Kemess South Project”, March 5.

     GRP (1997b). “Geotechnical Review Panel — Kemess Project”, June 6.

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     GRP (1997c). “Geotechnical Review Panel Report — Kemess Project”, October 21.

     GRP (1998a). “Commencement of Tailings Deposition, Kemess Mine Tailings Impoundment, Geotechnical Review Panel Comments”, May 12.

     GRP (1998b). “Kemess Mines Inc. — Independent Review Panel (IRP) Report”, June 18.

     GRP (2000). “Geotechnical Review Panel Report — Kemess Tailings Storage Facility”, June 27.

     GRP (2001). “Geotechnical Review Panel Report — Kemess Tailings Storage Facility”, October 2.

     “Geotechnical Investigations and Evaluation for Open Pit Development” (Ref. No. 11816/39-l), Dean R. Brox, P. Eng. and Ken J. Brouwer, P. Eng.

     “Report to Kemess Mine, Pit Slope Stability and Design,” C.O. Brawner, P. Eng.

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Appendix A    Assay Procedures

Assay Procedures for Copper and Gold

Copper (Cu) Assay Procedure

This digestion is used for assay level Copper analysis. The method has a detection limit of 0.001%. If the sample assay is greater than the highest calibrating standard then repeat then procedure using 0.200g ± 0.001g sample

Record keeping:

General:
A flask number must be clearly marked for each weighed item. Maximum batch size: 60

Duplicates:
Worksheets are printed with as maximum of 24 samples per sheet. A minimum of 10% the samples on each worksheet must be digested in duplicate. The standard convention for weighing duplicates is to weigh the lst, 10th and 20th sample on any worksheet. This will ensure that at least 10% of the samples are done in duplicate. If any of these three samples are not to be analyzed then an adjacent sample must be substituted in its place. The duplicates are weighed into flasks immediately following the last sample on the worksheet. The ordinal number(s) of the duplicate samples weighed must be written on the worksheet directly after the last sample on that sheet.

Standards:
The standards are weighed into flasks immediately following the last duplicate on the worksheet. The name and weight of the standards must be written on the worksheet directly after the last duplicate on that sheet. The reagent blank should always be first, followed by any standards. Currently we are using CanMet standard MP-la for Cu Assay

Combining Worksheets:
Assayers Canada carts can hold from 50 to 70 flasks. Two or more worksheets can be combined onto one batch provided that all of the required standards and duplicates can also fit on the same cart. The samples from each worksheet are weighed in sequential order with the duplicates for each worksheet immediately following the samples for that worksheet. All standards and blanks are weighed following the last sample on the last worksheet. A cart number must be clearly marked on each worksheet. Each worksheet on the cart must have a reference to the other worksheet that are weighed onto the same cart.

Reagents and supplies
HCl, cone, reagent grade, Anachemia 4955
HNO₃, cone, reagent grade, Anachemia 6525
HBr, cone, reagent grade, Anachemia 4950
Phosphoric acid assay flasks, 200ml, borosilicate glass, Anachemia 010-631
Rubber stoppers, Size 02
Utility Cart for flasks
DI Water

Procedure:

1) Ensure all record keeping has been completed. Weigh l.000+ 0.001g of sample into a 200ml volumetric flask.

2) Add 15 ml HNO₃

3) Wait until any strong reaction subsides, then add 5ml HBr. Care must be taken when adding both of these acids. Watch that the reaction inside the flask does not become so vigorous that the acid bubbles over the top. This is especially true for suspected high sulfide samples.

4) Wait until any strong reaction subsides. Reflux on a medium hotplate until all bromine gas has been evolved. (approx. 30 minutes)

5) Remove from hotplate and allow to cool. Add 25ml HCl

6) Wait until any strong reaction subsides, then return to hotplate. Take note of the reaction (as before) when returning flasks to hotplate ensuring that the samples do not boil over.

7) Digest until all soluble salts have been dissolved.

8) Remove from hotplate and allow to cool. Bulk to 200 ml with D.I. H₂O. Cap with rubber stoppers and mix thoroughly.

9) The samples are analyzed on the AA using the standard instrumental conditions.

Quality Control

Every batch of samples will contain at least 10% duplication and one Canmet certified standard.

Canmet standard data and acceptable ranges
MP-la (l/5)      0.288% ± 0.003%

Corrective Actions:

1.	Check all record keeping to ensure that no errors have been made in the analysis procedure.
2.	Examine AA for any physical reason for the failure. If an obvious problem is detected, re-analyze the existing digested sample solutions for the entire batch.
3.	If the above fails to produce passable data or if an obvious problem is not detected, re-digest and re-analyze the entire batch from the beginning. Be sure to double check all sample labeling and record keeping.

Gold (Au) Fire Assay Procedure

This procedure is used for assay level Gold analysis. The method has a detection limit of 0.01g/tonne. If the sample assay is greater than the highest calibrating standard then repeat then procedure using less sample. This procedure is offered only as a general outline for fire assay. Due to the varying sample matrices encountered, same parameters can be modified at the time of analysis by the fire assayer.

Record keeping:

General:
A pot, cupel and test tube number must be clearly marked for each weighed item. Maximum batch size: 30. Clear indications of samples transferred from Pot to Cupel to test tube must be recorded.

Duplicates:
Worksheets are printed with as maximum of 24 samples per sheet. A minimum of 10% the samples on each worksheet must be digested in duplicate. The standard convention for weighing duplicates is to weigh the 1st, 10th and 20th sample on any worksheet. This will ensure that at least 10% of the samples are done in duplicate. If any of these three samples are not to be analyzed then an adjacent sample must be substituted in its place. The duplicates are weighed into pots immediately following the last sample on the worksheet. The ordinal number(s) of the duplicate samples weighed must be written on the worksheet directly after the last sample on that sheet.

Standards:
The standards are weighed into pots immediately following the last duplicate on the worksheet. The name and weight of the standards must be written on the worksheet directly after the last duplicate on that sheet. The reagent blank should always be first, followed by any standards. Currently we are using two in-house standards for routine Au Assay: 96-3 and 97-3

Combining Worksheets:
Assayers Canada carts can hold from 50 to 70 pots. Two or more worksheets can be combined onto one batch provided that all of the required standards and duplicates can also fit on the same cart. The samples from each worksheet are weighed in sequential order with the duplicates for each worksheet immediately following the samples for that worksheet. All standards and blanks are weighed following the last sample on the last worksheet. A cart number must be clearly marked on each worksheet. Each worksheet on the cart must have a reference to the other worksheet that are weighed onto the same cart.

Gold (Au) Fire Assay Procedure

Reagents:
Pre-mixed Neutral Flux — Mines Assay Supplies

Approx:	Pb0	50	%
	Na2CO3	40	%
	Na2B407	7.5	%
	SiO2	2.5	%

1:3 Nitric acid
     - mix 1 part conc. Nitric acid to 3 parts D.I. water
HC1, conc, reagent grade, Anachemia 4955
Ag inquarts or AgNO3 solution
Flour as required
KNOs as required
D.I. Water

General Procedure:

1)	Weigh l/2 or 1 assay ton sample into 40g crucible
2)	Add 80, pre-mixed flux, 14g PbO and 5g Na2CO3. Mix well.
3)	Add 2.5mg Ag inquart
4)	Fuse at 1000 C for 1 hour
5)	Pour the melt into steel conical molds and allow to cool. Separate slag and lead button.
6)	Cupel lead button @ 925 C until complete (approx. 45 min.)
7)	Collect bead from cupel and place in disposable test tube. Add 2ml 1:3 HNO3 and part for l/2 hour in 70 C water bath
8)	Add 3ml cone. HCl and digest for 1/2 hour in water bath.
9)	Remove from bath and allow to cool. Dilute to 10ml with D.I. water and mix.
10)	Read on A.A. using appropriate calibrating standards.

Quality Control

Every batch of samples will contain at least 10% duplication and two in-house standards. All in-house standards must be verified by round robin analysis

Standard data and acceptable ranges
96-3    0.40g/tonne ± 0.03g/tonne
97-3    1.36g/tonne ± 0.l0g/tonne

Corrective Actions:

1.	Check all record keeping to ensure that no errors have been made.
2.	Due to the complex nature of fire assay, all failures of standard data results in a complete re-assay of the entire

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Appendix B    Sections

Section #10050
Section #10150

Section #10660 East
Section #9660 East