Exhibit 99.1

THE RED LAKE GOLD MINES PROPERTY

RED LAKE MINING DIVISION

NORTHERN ONTARIO

Prepared by:	Dean Crick, M.Sc., P.Geo.	November 17, 2006
	Stephane Blais, P.Eng.
	Anthony Stechishen, P.Geo.

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TABLE OF CONTENTS

SECTION			PAGE
TABLE OF CONTENTS			II
1.0	EXECUTIVE SUMMARY		1
2.0	GLOSSARY		3
3.0	INTRODUCTION AND TERMS OF REFERENCE		5
4.0	DISCLAIMER		5
5.0	PROPERTY DESCRIPTION AND LOCATION		6
	5.1	Location	6
	5.2	Property Status	6
	5.3	Reclamation and Permitting	6
6.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY		7
	6.1	Accessibility	7
	6.2	Climate	7
	6.3	Local Resources	7
	6.4	Infrastructure	8
	6.5	Physiography	8
7.0	HISTORY		8
	7.1	Red Lake Complex	12
	7.2	Campbell Complex	13
8.0	GEOLOGICAL SETTING		14
	8.1	Regional Geology	14
		8.1.1    Granitoid Plutons and Felsic Porphyry Intrusions	15
		8.1.2    Structure	15
		8.1.3    Metamorphism	16
	8.2	Local Geology	17
		8.2.1    Stratigraphy	17
		8.2.2    Structural geology	18
		8.2.3    Deformation History	18
		8.2.4    Metamorphism	19
	8.3	Site Geology	20
		8.3.1    Mineralogy	20
		8.3.2    Alteration	21
		8.3.3    Structure	21
9.0	DEPOSIT TYPES		22
10.0	MINERALISATION		25
	10.1	Planar to curviplanar zones	25
	10.2	Plunging zones	26

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	10.3	Other	27
	10.4	Zones	27
		10.4.1  Red Lake Complex Zones	27
		10.4.2  Campbell Complex Zones	28
11.0	EXPLORATION		31
	11.1	HGZ and FW Sulphides	32
	11.2	Deep Campbell	33
	11.3	Partywall Opportunities	34
	11.4	Surface Exploration Targets/ Concepts	34
	11.5	District exploration	35
	11.6	Additional Studies	35
12.0	DRILLING		35
	12.1	Red Lake Complex Diamond Drilling	38
		12.1.1  Drill Core Logging Procedures	38
		12.1.2  Mine Grid and Survey Control	39
	12.2	Campbell Complex Diamond Drilling	40
		12.2.1  Drill Core Logging Procedures	40
		12.2.2  Mine Grid and Survey Control	41
13.0	SAMPLING METHOD AND APPROACH		41
	13.1	Red Lake Gold Mines (Red Lake and Campbell Complexes)	41
	13.2	Chip Sampling	42
	13.3	Muck Sampling	42
	13.4	Testhole Sampling	43
14.0	SAMPLE PREPARATION, ANALYSES AND SECURITY		43
	14.1	Red Lake Complex	43
		14.1.1  Analysis (RLC)	44
		14.1.2  QA/QC Procedures (RLC)	44
	14.2	Campbell Complex	45
		14.2.1  Analysis (CC)	45
	14.3	QA/QC Procedures (CC)	46
15.0	DATA VERIFICATION		47
	15.1	Red Lake Complex	47
	15.2	Campbell Complex	48
16.0	ADJACENT PROPERTIES		48
17.0	MINERAL PROCESSING AND METALLURGICAL TESTING		48
	17.1	Red Lake Complex	48
		17.1.1  Red Lake Mill Performance and Limits	50
	17.2	Campbell Complex	54
		17.2.1  Crushing	54
		17.2.2  Grinding and Gravity Circuit	54

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		17.2.3 Flotation Circuit	55
		17.2.4 Pressure Oxidation Circuit	55
		17.2.5 Flotation Tails Leaching and Carbon-In-Pulp Circuit	57
		17.2.6 Refinery	57
		17.2.7 Paste fill and Waste Treatment Circuit	57
		17.2.8 Effluent Treatment Circuit	58
		17.2.9 Polishing Pond and wetland	58
		17.2.10            Planned Flowsheet Developments	58
		17.2.11            Production Performance	59
18.0	MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES		60
	18.1	Red Lake Complex	61
	18.2	Method of Resource evaluation (RLC)	63
		18.2.1 Database and Validation	63
		18.2.2 Geological Controls	64
		18.2.3 Zone Interpretation and Modelling	65
		18.2.4 Polygonal Mineral Resource Modelling	65
		18.2.5 Statistical Analyses	67
		18.2.6 Cutting and Capping Procedures	68
		18.2.7 Mineralized Intercepts	70
		18.2.8 Grade Estimation, Cutoff and Mine Operating Costs	71
	18.3	Application of Mineral Resource Definitions (RLC)	72
	18.4	Mineral Resource and Reserve Statement for High Grade Zone (RLC)	72
		18.4.1 Cut Grade Mineral Resource Estimate	73
		18.4.2 Uncut Grade of the Mineral Resource Estimate	73
		18.4.3 Mineral Resources by Zone	73
		18.4.4 Mineral Reserve Statement for High Grade Zone	75
		18.4.5 High Grade Zone Mineral Reserves	75
		18.4.6 Mineral Reserves By Zone	77
	18.5	Sulphide Zones Mineral Resource and Reserve Estimate (RLC)	79
		18.5.1 Application of Definitions for the Sulphide Zones	79
		18.5.2 Method of Estimation for Sulphide Zones	80
		18.5.3 Statistical Analyses	81
		18.5.4 Application of Mineral Reserve Definitions for the High Grade Zone	83
		18.5.5 Sulphide Zone Mineral Resources and Reserves	83
	18.6	Campbell Complex	87
		18.6.1 Mineral Resource Statement (CC)	87
		18.6.2 Mineral Reserve Statement (CC)	88
		18.6.3 Mineral Resource Methodology	89
		18.6.4 Mineralized Zone Model (CC)	90

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		18.6.5 Compositing and Capping	91
		18.6.6 Grade Estimation Methods	93
		18.6.7 Verification of resource estimation procedures - Model: 10_00_F (includes 3 zones)	94
		18.6.8 Verification of resource estimation procedures - Model: 308_27_56_561 (includes 4 zones)	95
		18.6.9 Verification of resource estimation procedures - Model: 36_30_mm_tp	96
		18.6.10            Verification of resource estimation procedures - Model: 42_36_DC	97
		18.6.11 Verification of resource estimation procedures - Model: 42_36_mm_tp_sxc (includes 3 zones)	97
19.0	OTHER RELEVANT DATA AND INFORMATION		110
	19.1	Production Plan and Schedule	110
		19.1.1 Mining Extraction and Ore Losses	111
	19.2	Red Lake Complex	112
		19.2.1 Ground Conditions	113
	19.3	Access Development	116
	19.4	Infrastructure	119
		19.4.1 Site Facilities	119
		19.4.2 Power	120
		19.4.3 Ventilation	120
		19.4.4 Dewatering	120
		19.4.5 Compressed Air	121
		19.4.6 Communication	121
	19.5	Campbell Complex	121
		19.5.1 Mine Plan	122
		19.5.2 Geology Summary and Ground Conditions	123
		19.5.3 Access Development	128
		19.5.4 Mining Method	128
		19.5.5 Equipment	129
		19.5.6 Mining Dilution	130
		19.5.7 Infrastructure	131
		19.5.8 Ventilation	131
		19.5.9 Dewatering	131
		19.5.10            Compressed Air	132
		19.5.11            Communication	132
	19.6	Environmental and Socio-economic	132
		19.6.1 Permitting Status	132
		19.6.2 Socio-economic and Environmental Conditions	134

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		19.6.3 Bonding, Reclamation and Closure	135
20.0	CERTIFICATE OF QUALIFICATIONS AND CONSENT OF QUALIFIED PERSONS		137
21.0	REFERENCES		138
22.0	ADDITIONAL FIGURES AND ILLUSTRATIONS		143

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LIST OF TABLES
Table 5.1	List of Claims registered for the Red Lake Mines
Table 9.1	Summary of geological characteristics of the Red Lake Mines and generic GQCV
Table 12.1	Reclassification of Lithological codes
Table 12.2	Location of Mine Surface Survey Points
Table 13.1	Comparison of 2005 Mining Reconciliation vs. Drilling, Chip and Muck Samples
Table 18.1	Red Lake Mine Mineral Reserves (as of December 31, 2005)
Table 18.2	Red Lake Mine remaining Mineral Resources (as of December 31, 2005)
Table 18.3	Basic Statistics on High grade Zone Raw Assays
Table 18.4	Statistics on High Grade Zone two-Foot (60cms) normalized assays
Table 18.5	Per Zone Top Cuts on High Grade Zone Two-Foot (60cms) normalized samples (December, 2005)
Table 18.6	Basic Statistics on High Grade Zone uncut mineralised intercepts
Table 18.7	Goldcorp Unit Operating Costs - 2005
Table 18.8	High Grade Zone Mineral Resources (Cut Grade) 2005
Table 18.9	High Grade Zone Mineral Resources (Uncut Grade) 2005
Table 18.10	High Grade Measured and Indicated Mineral Resources by Zone - 2005
Table 18.11	High Grade Zone Mineral Reserves (as of December 31, 2005)
Table 18.12	Total Production Compared to Estimated Mineral Reserves (to November 30, 2005)
Table 18.13	High Grade Zone Proven and Probable Mineral Reserves by Zone – 2005
Table 18.14	Tonnage and Grade Distribution of High Grade Zone Mineral Reserves - 2005
Table 18.15	Statistics on Sulphide Zones Two-Foot Normalized Uncut Assays (from 2004)
Table 18.16	Statistics on Two-Foot Normalised Uncut Assays on Selected Sulphide Zones for 2005*
Table 18.17	Basic Statistics on Sulphide Zones Raw Assays (from 2004)
Table 18.18	Basic Statistics on Raw Assays (Selected Sulphide Zones, 2005)
Table 18.19	Basic Statistics on Sulphide Zones Uncut Mineralised Intercepts (from 2004)
Table 18.20	Basic Statistics on Uncut Mineralised Intercepts (Selected Sulphide Zones, 2005*)
Table 18.21	Summary of Mineral Resources of the Sulphide Zones (as of December 31, 2005)
Table 18.22	Sulphide Measured and Indicated Mineral Resources by Zone - 2005
Table 18.23	Summary of Mineral Reserves of the Sulphide Zones (as of December 31, 2005)
Table 18.24	Sulphide Zones Proven and Probable Mineral Reserves by Zone Above Level 30 (as of December 31, 2005)

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LIST OF TABLES (CONT’D)
Table 18.25	Sulphide Zones Proven and Probable Mineral Reserves by Zone Below 30 Level (as of December 31, 2005)
Table 18.26	Mineral Resource Statement for Campbell Complex to September 2006
Table 18.27	Digs Resource Statement
Table 18.28	Mineral Reserve Statement for Campbell Complex
Table 18.29	Block Model Remaining Resources (Excluding Reserve)
Table 18.30	Statistics for Five Major Projects – Uncapped Composite Chip Data
Table 18.31	Statistics for Five Major Projects – Uncapped Composite Drillhole Data
Table 18.32	Capping Levels for Composited Assays for Five Major Projects
Table 18.33	Project 10_00_F - F Zone Significant Drill Intercepts
Table 18.34	Project 36_30_mm_tp - TP Zone Significant Drill Intercepts
Table 18.35	Project 42_36_mm_tp_sxc - TP Zone Significant Drill Intercepts
Table 18.36	Project 308_27_56_561 - 561 Zone Significant Drill intercepts
Table 18.37	Project 42_36_DC - DC Zone Significant Drill Intercepts
Table 19-1	HGZ Joint Sets
Table 19-2	RLC Mining Methods
Table 19.3	Mobile Equipment List for Red Lake Complex
Table 19.4	Material Properties of Rock Units
Table 19.5	Major Faults at Campbell Mine
Table 19.6	Joint Sets Identified at Campbell Mine
Table 19.7	In-situ Stress measurements for Campbell Mine
Table 19.8	Mine Wide Stress Tensor Campbell Complex
Table 19.9	Mobile Equipment at Campbell Mine Complex
Table 19.10	Environmental Permit and Approval List for Red Lake Gold Mine

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LIST OF FIGURES
Figure 5.1	Location Map for Red Lake Mines
Figure 5.2	Claim Map for Red Lake Mines
Figure 5.3	Location of Infrastructure for Red Lake Mines
Figure 7.1	Annual Production for Red Lake Complex
Figure 7.2	Resources and Reserves for Red Lake Complex
Figure 7.3	Annual Production for Campbell Complex
Figure 7.4	Resources and Reserves for Campbell Complex
Figure 8.1	Regional bedrock Geology Map for Red Lake Mines Area
Figure 8.2	Airborne Magnetics Map of the Red Lake Greenstone Belt
Figure 8.3	Local Geology Map for Red Lake Mines Area
Figure 8.4	Synoptic Model for the Structural Evolution of the Campbell_Red Lake System
Figure 8.5	Mine Site Geology for Red Lake Mines
Figure 10.1	Red Lake Gold Mines Ore Zones- 14 Level Plan
Figure 11.1	Combined Cumulative Production and Year-End Reserve Ounces
Figure 11.2	Deep Campbell Longitudinal Projection, showing Deep Campbell
Figure 11.3	Longitudinal Projection of Red Lake Gold Mines - Priority Exploration
Figure 11.4	Bounded Polygonal Longitudinal Vertical Section of The HGZ Below 37L
Figure 11.5	Longitudinal Projection of the FW Sulphides Between 30-40 Levels Showing Drill Target Balls for 2006
Figure 11.6	Longitudinal Projection of the Partywall Prospects
Figure 14.1	2005 Original assays versus Duplicate pulps for Red Lake Complex
Figure 14.2	Red Lake Complex Standards Assays
Figure 17.1	Process Circuit for Red Lake Complex
Figure 17.2	Annual Feed Ounces for Red Lake Complex
Figure 17.3	Annual Recovery Rates for Red Lake Complex
Figure 17.4	Comparison of Recovery rates against Feed Ounces for Red Lake Complex
Figure 17.5	Comparison of Grinding Tonnage to Recovery Rates for Red Lake Complex
Figure 17.6	Process Circuit for Campbell Complex
Figure 17.7	Monthly Gold Production for Campbell Complex
Figure 17.8	Monthly Mill Recovery for Campbell Complex
Figure 17.9	Monthly Mill Utility for Campbell Complex
Figure 17.10	Average daily Mill Tonnes by Month for Campbell Complex
Figure 18.1	Cross Section 6150SE Showing Typical Definition Drilling and Splay
Figure 18.2	Interpreted Geology of the 37 Level
Figure 18.3	HW5 – Vertical Longitudinal Projection Polygonal Resource Estimate
Figure 18.4	Main Zone – Vertical Longitudinal Projection Polygonal Resource Estimate
Figure 18.5	intentionally deleted
Figure 18.6	intentionally deleted

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LIST OF FIGURES (CONT’D)
Figure 18.7	Log Normal histogram of Selected Sulphides
Figure 18.8	Log Normal Probability plot for Selected Sulphides
Figure 18.9	10_00_F – F ZONE Probability plot uncapped 2 ft (0.6m) composited chips
Figure 18.10	10_00_F – F ZONE Probability plot uncapped 2 ft (0.6m) composited drill holes
Figure 18.11	10_00_F – F ZONE Histogram cap analysis for uncapped 2 ft (0.6m) composited chips
Figure 18.12	10_00_F – F ZONE Histogram cap analysis for uncapped 2 ft (0.6m) composited drillholes
Figure 18.13	308_27_56_561-561 ZONE Probability plot uncapped 2 ft (0.6m) composited chips
Figure 18.14	308_27_56_561-561 ZONE Probability plot uncapped 2 ft (0.6m) composited drill holes
Figure 18.15	308_27_56_561-561 ZONE Histogram cap analysis for uncapped 2 ft (0.6m) composited chips
Figure 18.16	308_27_56_561-561 ZONE Histogram cap analysis for uncapped 2 ft (0.6m) composited drillholes
Figure 18.17	36_30_mm_tp-TP ZONE Probability plot uncapped 2 ft (0.6m) composited chips
Figure 18.18	36_30_mm_tp-TP ZONE Probability plot uncapped 2 ft (0.6m) composited drill holes
Figure 18.19	36_30_mm_tp-TP ZONE Histogram cap analysis for uncapped 2 ft (0.6m) composited chips
Figure 18.20	36_30_mm_tp-TP ZONE Histogram cap analysis for uncapped 2 ft (0.6m) composited drillholes
Figure 18.21	42_36_MM_TP-SXC_TP ZONE Probability plot uncapped 2 ft (0.6m) composited chips
Figure 18.22	42_36_MM_TP-SXC_TP ZONE Probability plot uncapped 2 ft (0.6m) composited drill holes
Figure 18.23	42_36_MM_TP-SXC_TP ZONE Histogram cap analysis for uncapped 2 ft (0.6m) composited chips
Figure 18.24	42_36_MM_TP-SXC_TP ZONE Histogram cap analysis for uncapped 2 ft (0.6m) composited drillholes
Figure 18.25	42_36_DC – DC ZONE Probability plot uncapped 2 ft (0.6m) composited chips
Figure 18.26	42_36_DC – DC ZONE Probability plot uncapped 2 ft (0.6m) composited drill holes
Figure 18.27	42_36_DC – DC ZONE Histogram cap analysis for uncapped 2 ft (0.6m) composited chips

- xi - November 17, 2006

LIST OF FIGURES (CONT’D)
Figure 18.28	42_36_DC – DC ZONE Histogram cap analysis for uncapped 2 ft (0.6m) composited drillholes
Figure 18.29	42_36_DC – DC ZONE Probability Plot for 2 ft (0.6m) uncapped composited chips
Figure 18.30	42_36_DC – DC ZONE Probability Plot for uncapped 2 ft (0.6m) drillholes
Figure 18.31	42_36_DC – DC ZONE Histogram for uncapped 2 ft (0.6m) composited chips
Figure 18.32	42_36_DC – DC ZONE Histogram for uncapped 2 ft (0.6m) composited drillholes
Figure 19.1	Hoisting Capacity for Combined Operations
Figure 19.2	Milling Capacity for Combined Operations
Figure 19.3	Reserves and Production for Combined Operations
Figure 19.4	HGZ Principle Stresses
Figure 19.5	Campbell Complex Stress Measurements as a Function of Depth
Figure 19.6	Annual Reportable Ground Incidents for Campbell Complex

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1.0          EXECUTIVE SUMMARY

This technical report has been prepared for filing pursuant to National Instrument 43-101 (“NI 43-101”) Standards of Disclosure for Mineral Projects of the Canadian Securities Administrators in connection with mineral reserve and mineral resource (“MRMR”) estimates and certain other information relating to Goldcorp Inc.’s (“Goldcorp”) Red Lake Gold Mines (“RLGM”) as of September 30, 2006.  The format and content of this report are intended to conform to Form 43-101F1.

The Qualified Persons responsible for the preparation of this report are Dean Crick, M.Sc., P.Geo., Manager of Mine Geology, Goldcorp Canada Ltd. (RLGM), Stephane Blais, P.Eng., Chief Engineer, Goldcorp Canada Ltd. (RLGM), and Anthony Stechishen, P.Geo., Senior Resource Geologist, Goldcorp Canada Ltd. (RLGM).  This report has been prepared by employees of Goldcorp under the supervision of Messrs. Crick, Blais and Stechishen.  Information in this report is based on work conducted by Goldcorp geologists, engineers and metallurgists as well as third party consultants retained by Goldcorp.

A consolidated audit of the operations and MRMR of Goldcorp’s Red Lake Gold Mines, being the Campbell Complex (formerly operated by Placer Dome) and the Red Lake Mine, will subsequently be prepared for the 2006 Year End.

On May 12, 2006, Goldcorp Inc. acquired all of the issued and outstanding shares of Placer Dome (CLA) Limited, then a wholly owned subsidiary of Barrick Gold Corporation (formerly, Placer Dome Inc.) and operator of the Campbell Mine located adjacent to the town of Balmertown, Ontario which is 10 kilometres to the east of the community of Red Lake.  The name of Placer Dome (CLA) Limited was subsequently changed to Goldcorp Canada Ltd. by filing Articles of Amendment with Industry Canada and the Campbell Mine was integrated with Goldcorp Inc.’s Red Lake Mine, now the Red Lake Complex (“RLC”), to form one mining operation thereafter known as the Red Lake Gold Mines.  The Campbell Mine is now referred to as the Campbell Complex (“CC”) of RLGM.  The RLGM mine site is adjacent to the town of Balmertown, ten km east of the community of Red Lake.  Road access is by Highway 105, which joins the Trans-Canada Highway at Vermilion Bay, east of Kenora.

RLGM is an Archean lode gold deposit hosted along the eastern flank of the Red Lake greenstone belt, situated in the western portion of the Uchi Subprovince, a typical Archean granite-greenstone terrain containing eastward trending belts of volcanic and sedimentary rocks and syn-volcanic intrusives. The Red Lake camp is the second largest gold district in Canada with historical production in excess of 20 million ounces (1000 tonnes Gold).

- 2 - November 17, 2006

In general, there are three types of ore mineralization encountered at the Red Lake Gold Mines,, namely quartz-carbonate vein type ore, disseminated sulphide ore and auriferous silica (+/- sulphide) replacement ore.

Gold assays are conducted at the Campbell Complex assay lab (the “Campbell Lab”) and the SGS commercial lab located in Red Lake, Ontario. Process samples, mucks, chips and drill core are assayed using leading industry fire assay practices under the RLGM’s quality assurance and quality control (QA/QC) program. In addition to the Campbell Lab’s own QA/QC routine, the Red Lake Gold Mines geology group inserts various standard reference materials, blanks or duplicates roughly every twenty samples.

MRMR at RLGM are based upon geologically constrained models constructed by interpolating capped, composited assay values. This is where the similarities end, as grade estimation techniques applied at the Campbell Complex involve block models with ordinary kriging, inverse distance and nearest neighbour methodology versus polygonal longitudinal vertical sections at the Red lake Complex. WGM supports the 2005 MRMR at the Red Lake Complex, and recommended that inverse distance estimation be employed at the Campbell Complex for the restated MRMR dated to September 30, 2006. As an interim measure for this technical report, the MRMR for the Red Lake Complex is calculated as the 2005 audited figure minus the mined out to September 30, 2006. At the Campbell Complex the MRMR is re-calculated based on the WGM recommended reduction factor for the Mineral Resource and inverse distance estimation technique for the re-calculated Mineral Reserve. The Mineral Resources were re-estimated at a diluted cut-off grade ranging from 0.21-0.38 opt depending on specific ore lens diluted to a minimum horizontal width of 6.0 ft.

Actual production from January 2006 to September 30, 2006 is comprised of 143,116 ounces from the Campbell Complex and 353,220 ounces from the. Red Lake Complex.

Gold milling at both complexes includes a conventional crushing/rod and ball milling circuit for size reduction. On the processing side, gold is extracted by gravity, cyanide leach CIP (Carbon-In-Pulp) followed by electrowining and pressure oxidation autoclave recovery techniques for sulfide-rich concentrates. The Campbell Complex processing facility with a permitted capacity of roughly 1850-2000 tons per day, an average head grade between 0.30-0.4 opt and recoveries of 96% includes a pressure oxidation autoclave facility. The Red Lake Complex mill is operating at between 700-750 tons per day with a head grade of roughly 1.5- 2.5 opt, recoveries of 95-96%, and permitted capacity of 800 tons per day.

In 2006, a small open pit pilot project was completed at the Campbell Complex during the summer months producing roughly 45,000 tons of ore at an estimated grade of 0.10-0.12 opt. The open pit muck has been utilized to supplement the underground feed during periodic shortfalls.

- 3 - November 17, 2006

Current significant capital projects include the completion of the No. 3 Shaft at the Red Lake Complex to a vertical target depth of 6336 ft. and mill upgrade project to be completed by the end of 2007 at a remaining capital cost of $80 million.  This combined expansion project was designed to provide access to the HGZ via the internal ramp down to 47 Level (i.e. 7000 vertical ft), hoisting capacity of 6000 t/d and potential mill capacity upgrade to roughly 1000 t/d. An intermine connection between 34 level of the Red Lake Complex and 36 Level of the Campbell Complex has recently been completed allowing the potential transfer of muck from the Red Lake Complex across to the Campbell Complex Reid Shaft loading pocket to be hoisted to surface. The Life of Mine (LOM) sustainable capital upon completion of the expansion project is estimated to be roughly $25-$30 million annually.

The effective date of this technical report is November 17, 2006. The following are the Qualified Persons (“QPs”) for this technical report are all employed at the RLGM and the section(s) for which they are responsible set out next to their respective names.

QP	Sections
Dean Crick, M.Sc., P.Geo.	1, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, 18
Anthony Stechishen, P.Geo.	12, 13, 14, 15, 16, 18
Stephane Blais, P.Eng.	17, 18, 19

2.0          GLOSSARY

Centimetre	cm
Cubic centimetre	cm3
Cubic metre	m3
Day	d
Days per week	d/wk
Days per year (annum)	d/a
Degree	°
Degrees Celsius	°C
Gram	g
Grams per tonne	g/t
Greater than	>
Hectare (10,000 m2)	ha
Hour	h
Hours per day	h/d
Kilogram	kg
Kilograms per cubic metre	kg/m3
Kilograms per hour	kg/h

- 4 - November 17, 2006

Kilograms per square metre	kg/m2
Kilometre	km
Kilometres per hour	km/h
Less than	<
Litre	L
Metre	m
Metres above sea level	masl
Metres per minute	m/min
Metres per second	m/s
Micrometre (micron)	µm
Millimetre	mm
Million	M
Million tonnes	Mt
Minute (plane angle)	‘
Minute (time)	min
Ounce (troy ounce - 31.1035 gms)	oz
Percent	%
Pound(s)	lb
Parts per million	ppm
Parts per billion	ppb
Second (plane angle)	“
Second (time)	s
Square centimetre	cm2
Square kilometre	km2
Square metre	m2
Short Tons (907 kgs)	tons
Tonnes (1000 kgs)	t
Tons (short) per day	tpd
Tonnes per day	t/d
Tonnes per hour	t/h
Tonnes per year	t/a
Year (annum)	a
* Note all conversions from imperial to metric have been rounded

- 5 - November 17, 2006

3.0          INTRODUCTION AND TERMS OF REFERENCE

As a component of the Barrick Gold Corporation’s acquisition of the Placer Dome Group initiated in October 2005, Goldcorp Inc. acquired Placer Dome’s interests in the Porcupine Joint Venture, Musselwhite and the Campbell Mine located in Ontario by acquiring the operator of those mines, Placer Dome (CLA) Limited (now Goldcorp Canada Ltd.).

The acquisition closed on May 12, 2006 and the two formerly independent and adjoining Red Lake Mines and the Campbell Mines were integrated into a single operation to form the new Red Lake Gold Mines.  After a long history spanning seven decades of the two operations mining separate portions of the same orebody, they were finally merged into a single operation.

The 2005 Year End MRMR estimates for the Red Lake Complex were calculated in January, 2006 and have since been verified by WGM personnel, with the audit document authored by Michael Kociumbus, P.Geo Ross McFarland, P.Eng and Bruce Brady, P.Eng. Robert de l’Etoile of WGM has twice visited the Campbell Complex to confirm that the recommendations made by WGM in respect of the Mineral Resource estimates methodology at Campbell Complex were applied in the preparation of this report.

The majority of the information contained within this report has been supplied by the RLGM personnel.  The Mineral Resource estimates were prepared under the guidance of Dean Crick, P.Geo. (Mine Geology Manager, RLGM), Anthony Stechishen, P.Geo. (Senior Resource Geologist, Campbell Complex) and Mark Epp (Senior Resource Geologist, Red Lake Complex). Mineral Reserves were prepared under the guidance of Stephane Blais, P.Eng. (Chief Engineer, RLGM).  Mssrs. Blais, Crick and Stechishen, each of whom is being designated as Qualified Persons (“QPs”) with the ability and authority to verify the authenticity and validity of this data, are responsible for the preparation of this NI 43-101 compliant technical report. This document summarizes the professional opinion of the QPs and includes conclusions and estimates that have been based on professional judgement and reasonable care. All conclusions and estimates are based on the assumptions and conditions outlined in this report as of November 17, 2006. The report is to be issued and read in its entirety. Written or verbal excerpts from this report may not be used without the expressed written consent of the QPs or senior management of Goldcorp Inc.

All references to dollars within this report are in Canadian Dollars.  For an explanation of abbreviations please refer to the Glossary in Section 2.0.

4.0          DISCLAIMER

Much of the Red Lake Complex information in this report was taken from “An Audit of the Mineral Reserves and Resources of the Red Lake Mine as of December, 2005 for Goldcorp Inc”,

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dated May 2006 and authored by Michael Kociumbas, B.Sc., P.Geo., Bruce Brady, B.Eng., P.Eng. and G. Ross MacFarlane, P.Eng.

5.0          PROPERTY DESCRIPTION AND LOCATION

5.1          Location

Goldcorp’s Red Lake mines cover approximately 2335 hectares located in the Red Lake Mining Division within the District of Kenora, northwestern Ontario (claim sheet G-3735). They lie within the Canadian Shield, approximately 180 kilometres north of Dryden, on Highway 17 (Figure 5.1). The properties lie within NTS map-sheet 052N04 at Latitude 51o 05’ 58” and Longitude 93o 43’21”W, UTM (NAD 27) coordinates 5653000N and 445400E, Zone 15 and are about 120 kms east of the provincial border with Manitoba.

5.2          Property Status

The Red Lake Complex consists of 89 patented mineral claims covering 1254 ha and the Campbell Complex consists of 77 patented mineral claims covering 1,084 ha.

Table 5.1 in Appendix A lists the patented mining claims and licences of occupation comprising the Red Gold Lake Mines.

Figure 5.2 shows the location of claims with the property outline.  The property has been legally surveyed for all the patented mining lands.

Figure 5.3 shows the location of the RLGM infrastructure with relation to the boundaries of the properties with an outline of the footprint of the sub-surface development, projected to surface.  Red Lake Gold Mines are serviced by four operating shafts, including an internal shaft from for the haulage of men, equipment, materials and rock.  There is also several sub-surface ramps, providing material and equipment movement to mining horizons at both complexes.  The Red Complex is nearing completion of their No. 3 Shaft which lies 3,800 ft (1,150m) south of the RLC No. 1 Shaft.

5.3          Reclamation and Permitting

The Red Lake Complex has a well-established environmental program.  Federal and provincial environmental legislation is recognized through policies and practices on site that ensure compliance.  The most comprehensive on-site environmental program covers the tailings effluent monitoring.  A mine closure plan has been submitted to the Ministry of Northern Development

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and Mines and the technical portion of the document has been accepted.  The tailings ponds are situated on the surface rights in the Balmer Township.

The Campbell Complex dealt with the negative environmental consequences of arsenopyrite contamination in waste rock and tailings pond discharge by constructing a nine hectare wetland. Copper, ammonia and low levels of other contaminants were removed through naturally occurring water treatment during the ice-free season. This process made Campbell one of the first gold mines in Ontario to harness natural, microbial forces to produce a non-toxic discharge.  This system is being introduced to the Red Lake Complex.

Of the Campbell Complex’s 160 hectares of land comprising its tailings management facility, 68 hectares are being used and the rest are at some stage of reclamation.

The tailings management facility is currently undergoing modifications to address current mill and mine effluent to ensure discharge water continues to meet all provincial and federal regulations.

For additional information see Section 19.5 of this report.

6.0          ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

6.1          Accessibility

The Red Lake area is accessible by Highway 105, which joins the Trans Canada Highway at Vermilion Bay, 175 kilometres to the south and 100 kms east of Kenora. Commercial air services from Thunder Bay and Winnipeg are available with several flights daily to the community.

6.2          Climate

The climate of the Red Lake area is typical northern continental boreal climate, with warm summers and cold winters. Temperatures ranges are 18 to 25+ degrees Celsius in July to -20 to -35 degrees Celsius in January. Annual precipitation is 310 mm.  In this township, snow usually starts falling around late October / early November, and starts melting around March but doesn’t fully melt until late April.

6.3          Local Resources

The Red Lake Gold Mines are located in northwestern Ontario, Canada.  The Red Lake Municipality is comprised of five towns, Madsen and McKenzie Island (outlying communities),

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Red Lake, Cochenour, and Balmertown, the home of the Red Lake and Campbell Complexes. The Municipality of Red Lake has a population of 5,500.  Other industries in the area are logging and tourism.

6.4          Infrastructure

Local infrastructure is well developed in the Red Lake area, with all the amenities needed by both the Red Lake and Campbell mines and the community available locally.

Potable water is supplied by the municipality. Process water is taken from Balmer Lake and Sandy Bay. Ontario Hydro supplies power. Each mine has 30 megawatts of transformer capability but is limited to 26 megawatts by the incoming line. There are multiple transformers throughout the sites, which step voltage down to 4160, 2300 and 575 volts.

6.5          Physiography

Topography is composed of typical Canadian Shield with irregular hummocky hills and discontinuous ridges created by glaciofluvial material and till.  These are separated by depressions and hollows occupied by lakes, ponds and muskeg. Much of the Red Lake region is still untouched and is accessible only by air or canoe.  The water level of Red Lake lies at 354 meters above sea level (masl) and elevations rise up to 50m above this.

The natural vegetation is predominantly black spruce, fir, larch (tamarack) and pine.  In the better drained elevated areas this changes to poplar, birch, willow, alder and mountain ash, with a variety of shrubs.

Bedrock outcrops are scattered and consist of less than 5% of the surface area.  Soil in the vicinity of the Red Lake and Campbell mines is characterized by a 30 to 50cm layer of topsoil overlying compact sand with traces of clay, gravel and scattered cobbles and boulders.  Low-lying areas contain silty-clay sediments that were deposited in glacial lakes.

7.0          HISTORY

The first Red Lake Gold Rush, which began in 1926, resulted in the discovery and development of several reasonably successful, former producing mining operations (i.e. >10 years production, >200,000 oz produced) including Cochenour/Willans, Madsen, Howey Bay Gold Mines, Mckenzie Red Lake, and Hasaga. The second gold rush which began in 1944 on the south shore of Balmer Lake where the town of Balmertown exists today resulted in the discovery of two prolific gold mines, the Dickenson Mine and Campbell Mine.  In operation since 1948 and 1949

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respectively, the following is a timeline from discovery to the present of significant highlights for the recently integrated Red Gold Lake Mines.

1944	January 12-14, 1944. George and cousin Colin Campbell restake lapsed 12 claim units south shore of Balmer Lake with partner, A.K. Mcleod.
1944	July 18, 1944. Partners registers as a limited liability company, The Campbell Red Lake Mines Ltd. with the assistance of mining promoters and underwriters Arthur W. White and John M. Brewis of the firm Brewis & White Ltd. who became executive officers in the company and helped underwrite the initial share offerings of the company.
1944	Oct 4, 1944 George Campbell blasts trench on the southern shore of Balmer Lake along a sheared mafic felsic contact returning > 1 opt grab samples and advertises preliminary assay results in Northern Miner.
1944	James McCrea, VP Dome Exploration, a wholly owned subsidiary of Dome Mines Limited, convinces corporation to option the property and subsequently take a controlling interest in Campbell Red Lake Mines Limited.
1945	Dickenson Red Lake Mines Limited was incorporated by the same promoters Arthur W. White and John M Brewis, who purchased the adjoining land packange from Toronto Red Lake Mines. After unsuccessful attempts at farming out the Dickenson Mine, Arthur W. White decides to try his hand at operating the mine.
1945	February 8, 1945 Discovery hole, the eighth hole in a stratigraphic fence hits significant visible gold in vein structure, just north of the former Campbell Mine administrative building.
1946	The head frame at Dickenson Red Lake Mines Ltd. was completed and shaft sinking had started. By the end of 1946, 2,667 metres of cross cuts, drifts and raises had been to a depth of 166 metres. In 1947, Highway 105 was completed, linking Red Lake with the Trans-Canada Highway.
1949	Underground mine production started at Campbell Mine.
1949	New Dickenson Mines Ltd. succeeds Dickenson Red Lakes Mine Ltd.
1953	New Dickenson Mines Ltd. acquires 16 claim Detta property from Detta Minerals Ltd.

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1959	Campbell Mine invests in consortium developing HG Young Mine, south shore Balmer Lake sinks shaft to 1000 feet dept, produces 59,000 oz Au until closing in 1962.
1960	Merger between New Dickenson Mines Ltd and Lake Cinch Mine to from Dickenson Mines Ltd, interest in Craibbe- Fletcher Gold Mines acquired.
1960	Campbell Mine options Craibbe-Fletcher from Dickenson Mines ltd.
1962	Drift from Dickenson Mine enters east adjoining Robin Red Lake Mines Ltd. ground between 18-23 Levels.
1963	Dickenson No.1 Shaft completed to 3589 ft., plan winze No.2 shaft 2960 ft SW, from 23rd to 30 Level, 3365 ft to 4838 ft between 1966-68.
1970	Dickenson Mines Ltd. acquires 80.3% of Robin Red Lake Mines Ltd. to begin production.
1975	Campbell Mine acquires control of HG Young claims
1978	Amalgamation of Dickenson Mines Ltd. and Robin Red Lake Mines Ltd.
1978	Dickenson no.2 shaft deepened to 5600 ft.
1982	Dickenson renamed Arthur W. White Mine; increased mill capacity to 700 tpd, sold 35% interest to Sullivan Mining group for $11M.
1983	Goldcorp launched as a separate holding company for gold mining shares and bullion.
1987	Merger between Campbell, Dome and Placer Development created Placer Dome Inc.
1989	In April 1989 Goldcorp obtained control of the Dickenson Mine at Red Lake.
1991	Autoclaving replaced roasting in the mill at Campbell.

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1995	The reorganized Dickenson operation invests $10 million in an exploration program which lead to the discovery of a high grade zone.
1995	Reid Shaft feasibility from surface approved at Campbell Mine.
1995	Significant discovery announced at Red Lake Mine – nine holes averaging 311.31 grams of gold per tonne across 2.3 metres.
1996	Craibbe Fletcher Mines Ltd. Gold Mines acquired by Placer Dome Canada Ltd.
1996	Decline to Levels 27 through 30 at Campbell was 50% complete by July, allowing for extraction of first development ore from depth development project.
1996	June 1996, the unionized work force at the Dickenson Red Lake Mine walked off the job, and a 46 month strike ensued.
1997	Mill flotation circuit upgraded at Campbell.
1998	Placer absorbs Lassie Red Lake Mines property adjacent Campbell Mine
1998	Construction of paste fill plant started at Campbell.
1999	Operation of the Reid Shaft at Campbell commenced and the paste fill plant was commissioned.
2000	Goldcorp settled deal with the union to walk away, and when production restarted the focus was on the HGZ below 30 Level
2000	Exploration development began on 39 level from the Reid Shaft at Campbell.
2000	Goldcorp posted all its proprietary geological data on the Web, Goldcorp Challenge - and gave away US$575,000 in prize money to geologists who proposed the best strategy for finding Red Lake’s next six million ounces of gold, above and beyond the company’s own results.

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2001	Discovery on the DC Zone at Campbell. Depth development project was completed. Campbell shaft (#1) was converted to people and material movement only. All hoisting (ore/waste) is skipped from the Reid shaft.
2002	Feasibility completed on the DC Zone. Pre-production development commenced at Campbell.
2003	Construction of the DC zone infrastructure progressed. Ore production started ahead of schedule at Campbell.
2005	Goldcorp completed its merger with Wheaton River Minerals.
2005	Intersection of the Red Lake mine’s “high grade zone” at depth (8,500 ft/2,600m) on the Campbell Mine Property.
2006	Goldcorp acquires Campbell mine to create the combined Red Lake Gold Mines.
2006	Red Lake Gold Mines completes a fibre optic network and inter-mine road which provides the backbone for the new mine’s communication and transportation system and connects surface to underground for video, voice and data.
2006	No. 3 Shaft excavation at Red Lake Complex completed to vertical depth of 6,336 ft (1,931m).

7.1          Red Lake Complex

Goldcorp Inc. was the continuing corporation formed on March 31, 1994 by the amalgamation of Goldcorp Inc., Dickenson Mines Limited, Goldquest Exploration Inc. and CSA Management Limited.  The company’s Red Lake gold mine is now known as the Red Lake Complex.

The Red Lake property was first staked during the Red Lake Gold Rush in 1926.  In 1944, the property was restaked and Dickenson Red Lake Mines Limited was incorporated.  Production mining began in 1948 at a rate of 125 tons per day (“tpd”) and increased to 500 tpd in the 1970s, with grades varying between 0.45 and 0.61 oz Au/ton.  In the early 1980s, the mill capacity was increased to 1,000 tpd and long-hole stoping was introduced which resulted in a severe drop in production grade. Cut-and-fill mining was subsequently re-introduced and production was increased to about 1,000 tpd by 1993/1994 with the grade ranging from 0.25 and 0.34 oz Au/ton.  The average recovery rate was 82%. Figure 7.1 illustrates historical annual production and annual

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reserves from the Red Lake Complex, formerly the Red Lake Mine and the Dickenson Mine since the mine first opened in 1948.

An exploration diamond drilling program initiated in 1995 within the lower levels of the mine resulted in the discovery of a cluster of High Grade gold veins between the 30 and 39 levels of the mine which are known collectively to as the High Grade Zone.

Between June 1996 and late 1999, operations were suspended due to a strike.  Mine staff and outside contractors maintained essential services and supported the exploration program on the property.

In September 1998, the feasibility of mining the High Grade Zone Mineral Reserves through a combination of existing mine infrastructure, new development, and a new processing facility purchased from Cameco (the Contact Lake mill), was studied.  An operating plan was developed to mine the Mineral Reserves by the Mechanized Cut-and-Fill Stoping Method at a rate of 600 tpd or 219,000 tons per year (tpy).  Using the Feasibility Study Mineral Reserve base, a mine life of 6.5 years was projected.

Mining from the High Grade Zone began in early 2000 and test milling in July.  By year-end, full rated production was achieved with a small stockpile of several thousand tons of crushed ore remaining on surface. Since production restarted in 2000 the exploration to increase resources has focussed on the High Grade Zone (HGZ).  Figure 7.2 illustrates that although the average tons of reserves has remained fairly consistent over the last 20 years, due to the considerably higher grades of the HGZ the amount of ounces of gold in the resources and reserves has increased substantially.

Since the beginning of operations in 1948 until the end of 2005 9.82 Million tons at 0.61 opt (5.99 M oz) has been processed from the Red Lake Complex.

7.2          Campbell Complex

The first recorded prospecting in the Red Lake District was carried out by the Northwestern Ontario Exploration Company in 1887, but gold was not discovered here until 1922 and the Campbell claims were not staked until 1926.  Then followed a period of claim cancellations and restaking of this area leading up to the cousins George and Colin Campbell restaking it in 1944.  At that time Campbell Red Lake Mines was incorporated and an option sold to Dome Mines, that they would take up to eventually lead to 57 percent ownership of the company.

In 1946 after further exploration of the property had been carried out, a four-compartment shaft was sunk to a depth of 597 ft (182m) with four levels leading off it.  In 1948 began the

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construction of the mill and the completion of the all weather road joining Balmertown (and the mine) to the main road leading from the Trans Canada Highway into Red Lake.  The mill went into operation the following year and soon reached its 300 tons per day capacity.

By 1953 the shaft had been deepened further to 2150 ft (655m) and soon after a deeper higher grade zone was discovered at the 14^th level.  For the following 30 years production at Campbell Mine remained consistent. Following the merger between Campbell, Dome and Placer Development to form Placer Dome Inc. a new phase of investment in the mine was begun.  This led to autoclaving replacing the roasting in the mill at Campbell; a decline from Levels 27 to 30; a mill flotation circuit upgrade; construction of paste fill plant and the commissioning of the Reid Shaft in 1999.  The latter enabled the discovery and development of the DC Zone.

Figure 7.3 and Figure 7.4 show the annual production from the Campbell Complex since the mine first opened in 1949. Since the beginning of mining operations in 1949 to the end of 2005, 19.17 Million tons at a grade of 0.58 opt (11.2 M oz) has been produced from the Campbell Complex.

8.0          GEOLOGICAL SETTING

8.1          Regional Geology

The Red Lake greenstone belt is situated in the western portion of the Uchi Subprovince, a typical Achaean granite-greenstone terrain containing eastward trending belts of volcanic and sedimentary rocks and syn-volcanic intrusives.  The rocks consist of volcanic and sedimentary assemblages representing magmatic and erosional events occurring over a period of approximately 300 million years.

The Red Lake greenstone belt outcropping is defined by an east-northeast orientated bow tie shaped, anticline, roughly forty kilometres in its long dimension and twenty kilometres across. Two perpendicular axes of mirror symmetry trend east-northeast and southwest defining the bow tie and intersecting at the Dome Stock near Red Lake. The most prominent axis trends east-northeast defining a broad antiform with older rocks of the Balmer Assemblage occupying the hinge flanked by younger rocks of the Huston and Confederation assemblages on its northern and southern flanks. Between the oldest and youngest assemblages is a regional unconformity, overlain by the Huston conglomerate.  At each end of the anticline, late-orogenic plutons stope up into the greenstone; the Killala-Baird batholith (2704 Ma) in the west., and the Walsh Lake pluton (2699 Ma) and Cat Island pluton (2697 Ma) in the east.

The belt is subdivided into several distinct assemblages. They include (from oldest to youngest) the Balmer assemblage (2.99-2.96 Ga), predominantly tholeiitc and komatiitic mafic to ultramafic extrusive rocks, the Ball Assemblage (2.94-2.92 Ga) calc-alkalic volcanic rocks in the

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northwest portion of the belt. The Balmer assemblage hosts the RLGM and constitutes over 50% of the Red Lake greenstone belt rocks. The sequence is disconformably overlain by a volcaniclastic/ clastic sequence, the Bruce Channel assemblage (2.89 Ga) prominent in the east and the Slate Bay assemblage (<2.92 Ga). In the southwest the Trout Bay assemblage (2.85 Ga) is characterized by predominantly mafic tholeiitic extrusive rocks. A laterally extensive polymict conglomerate belonging to the Houston assemblage marks an angular unconformity between Mesoarchean and Neoarchean strata. and is overlain by dominantly calc-alkalic felsic volcanic rocks belonging to the Confederation assemblage (2.75-2.73 Ga). These assemblages were deformed coeval with emplacement of the main phases of granitoid batholiths bordering the belt, followed later by the intrusion of the discordant granitoid Dome Stock.  Most of the supracrustal rocks have undergone only low-grade metamorphism. All the gold production in the red lake camp has come from the eastern half of the belt (Figure 8.1).

8.1.1       Granitoid Plutons and Felsic Porphyry Intrusions

The Red Lake greenstone belt is surrounded by several granitic batholiths. The McKenzie and Dome stocks intrude the centre of the Red Lake belt and host several past producing gold mines. The Howey diorite is an elongated intrusive stock measuring 3 miles by 1 mile, ranging in composition from gabbro to quartz diorite. It is in turn cut by a quartz porphyry dyke, which hosts the Howey and Hasaga ore bodies. Other small felsic intrusions hosting disseminated and vein mineralisation within zones of hydrothermal alteration are the Wilmar granodiorite, the Abino granodiorite and the Red Crest stock.

8.1.2       Structure

Large scale folding and steep fault systems dominate the eastern part of the Red Lake greenstone belt. Northeast trending synclines and anticlines occur northeast of McKenzie Island to the west of East Bay. On the eastern side of East Bay, a major fold with a northwest trending fold axis occurs in Bateman and Balmer Townships. In the Red Lake Gold Mines system, horizons of ultramafics within the basalt-dominated stratigraphy created complex geometries through competency contrast during the mechanical interaction of folding. High strain corridors characterized by pervasive foliation and cleavage development along mafic, ultramafic contacts are clearly distinguishable features using surface mapping and geophysics. Figure 8.2 superimposes mapped deformation corridors outlined in the mid to late 1980’s when the shear-hosted gold model became popular on the airborne magnetics for the Red Lake Greenstone belt.

Deformation prior to the deposition of the Bruce Channel assemblage resulted in locally angular relations of the underlying Balmer Assemblage with the Bruce Channel unconformity. Postdating the Bruce Channel and probably also the Confederation assemblages, D1 deformation folded the rock mass about N-S to NE-trending F1 fold axes. These folds apparently exhibited steep axial

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surfaces and may have formed through diapiric intrusive triggered gravity sliding or thrust repetitions of some packages.

Northeast-southwest (NE-SW) directed D2 compression resulted in F2 folding of the Balmer assemblage mafic/ultramafic rocks about steeply plunging axes in regions where stratigraphy was already steeply dipping as a result of F1 and earlier deformation. These folds, which are regional F2 structures, are highly non-cylindrical, and vary from West to East plunging with the majority around the Red Lake Gold Mines plunging moderately to the west. Associated with these D2 fabrics and folds are a family of south dipping, Northwest to East-West( E-W) striking D2 shear zones which transect the region. They are approximately axial planar to the F2 folds and likely represent the dominant regional S2 foliation trending 110-120° azimuth and dipping 65-85° to the southwest developed throughout the eastern portion of the belt.

Quartz-carbonate veining and alteration began developing at this stage focused around mafic-ultramafic contacts, particularly fold hinges and in the damage zones of small and large displacement fault/shear zones. Continued shortening and foliation development caused rotation of conjugate shear zones towards parallelism with F2 fold axial surface orientations, producing strong deformation and strain partitioning.  Further shortening resulted in strong near-vertical extension and locally the development of reverse faults.

Auriferous, siliceous, sulphidic alteration and quartz veining, overprinting earlier quartz-carbonate veins, began to develop at this stage, apparently after the bulk of strain and displacements on major fault systems. Minor strike slip movements on fault systems, reflecting at least local changes in the orientation of stress axes, contributed to the dilation and replacement, particularly in ultramafic/ mafic defined fold hinge zones abutting fault system.

Significant lateral and vertical displacements on shallower dipping fault systems may have occurred postdating mineralisation. These movements may largely reflect reactivation of the main fault network.

Mafic and porphyry dykes were intruded during late-stage, weak deformation largely following fault systems. Minor fault movements and late vein formation continued into the retrograde metamorphic history of the Red Lake Gold Mines system.

8.1.3       Metamorphism

Owing to the volume of late tectonic granitic plutons, a complex history of metamorphic events exists which arises from episodes of burial, crustal thickening, thrusting, baking by felsic intrusions of all sizes and ages, and late hydrothermal alterations focussed along shear zones.

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Within the Red Lake belt the alteration is characterized by greenschist metamorphic mineral assemblages, but with amphibolite facies mineral assemblages in areas close to the major plutons.

8.2          Local Geology

The gold bearing zones of the Red Lake Gold Mines occur on the eastern flank of a D2, high strain corridor, known locally as the Cochenour-Gullrock Lake deformation zone or Red Lake Mine trend.  These units are subparallel to a regional foliation that strikes from 100° to 120° azimuth, dipping southwest from 65° to 80° (Figure 8.2).

8.2.1       Stratigraphy

The Bruce Channel assemblage rests on the Balmer assemblage, with angular unconformity in some areas, to the southwest, north and east of the Red Lake Gold Mines.  The unconformity is inferred locally from the persistent along strike occurrence of the Bruce Channel basal felsic volcanics, which is concordant with the enveloping surface bounding the Balmer assemblage.  Where this felsic unit is lacking, and where the overlying clastics and iron formations of the Bruce Channel are in direct contact with Balmer assemblage, fault contacts can be inferred.  (Figure 8.3)

Ultramafics have been mapped from magnetics and drilling in the southern bordering, Lassie and Craibbe-Fletcher properties. One of these, which strikes west-northwest (WNW) through the boundaries of the properties is apparently thick and continuous, but offset by a WNW-trending fault into two bodies The weaker magnetic signature of these bodies toward the southeast, is indicative of strong alteration. Large mafic/ultramafic bodies, with relatively strong magnetic signatures have been mapped from the surface extension of the South Cross-Cut fault trending N to NNW until apparently truncated by the Bruce Channel contact (Figure 8.3).

Iron formations (strongly magnetic) and other units in the Bruce Channel outline a classic refolded synformal pattern, clearly recognizable from magnetics, to the north of Campbell and east of H G Young (Figure 8.3).

To the north and west of the Red Lake Complex the broad extent of Balmer Assemblage metabasalts are compatible with shallowly dipping enveloping surfaces. The magnetic signature in these areas, the occurrence of variable basalts and rhyolite packages containing iron formations (probably correlatable with those in the Red Lake mine corridor), may indicate that these are underlain by ultramafics, but at substantial depth. These ultramafics reach the erosion surface in the form of the East Bay Serpentinite.

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The thin east-southeast (ESE) striking magnetically anomalous zone to the north of the Bruce Channel synform hinge zone, apparently contains mafics, ultramafics and metasediments. A large irregular magnetic anomaly on the west of this zone may represent a fold interference pattern involving mafic/ultramafic intrusions.

8.2.2       Structural geology

NW-trending F2 folds appear to have been superposed on the North-South (N-S) to North-Northeast (NNE) trending axes of an earlier fold (and possibly thrust) system which appears to have had steep axial surfaces and shallow plunging axes. Major plunge variations of F2 folds appear to relate to their superposition on variably dipping stratigraphy, reflecting D1 folding. The F2 plunges in the Red Lake Gold Mines property are mostly steep, apparently due to a zone F1 folds through this region.  The F2 plunges to the north of the major synform creating Bruce Channel east of H G Young are commonly relatively shallow.  The F2 folds west of the Campbell properties exhibit shallow-to-moderate plunges.

At broad scales the Red Lake Gold Mines system appears to be hosted by a steeply plunging reclined F2 fold system with South-Southwest (SSW)-dipping axial surface, outlined by the Bruce Channel contact to the north, south and east, and cored by the Balmer mafic-ultramafic assemblage. This fold system has undergone significant modification by a system of steeply-to-moderately southwest dipping fault zones which are interpreted to have developed during D2 deformation, a NE-SW oriented shortening.

8.2.3       Deformation History

The D2 event comprised of progressive northeast southwest shortening resulted in the dismemberment of the F2 folds along axial planar shear structures, displacing the steeply plunging parasitic mafic/ultramafic fold hinges into discrete litho-structural domains.

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Figure 8.4 Synoptic model for the structural evolution of the Red Lake Gold Mines system.

8.2.4       Metamorphism

A recent assessment of metamorphic assemblages across the Red Lake greenstone belt place the Red Lake Gold Mines deposit on the low-grade side of the lower boundary of the transition zone between greenschist and amphibolite facies, and less than 200m from the biotite isograd marking the lower/upper greenschist-zone boundary.

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8.3          Site Geology

A comprehensive summary of the geologic setting of the Red Lake Gold Mines deposit is presented below. Readers should refer to the following papers for more details: MacGeehan and Hodgson (1982), Andrews et al. (1986), Rogers (1992), Penczak and Mason (1997, 1999), Zhang et al. (1997), Tarnocai (2000), Sanborn-Barrie et al. (2001, 2004), Twomey and McGibbon (2001) and Dube et al. (2001, 2002, 2003, 2004).

The Red Lake Gold Mines property is underlain mainly by tholeiitic basalt and locally by komatiitic basalt of the Balmer assemblage. The mine sequence is completed by peridotitic komatiite, rhyolite and associated mafic intrusions of the Balmer assemblage (figure 8.5). The steeply plunging south, southwest folded package is unconformable overlain by felsic volcaniclastics, clastic and chemical sedimentary rocks of the Bruce Channel assemblage defining an enveloping syncline anticline couplet based on younging directions, with the synform hinge located on the northside of the Campbell Complex, east of the HG Young shaft underneath Balmer Lake and the anticlinal hinge in the south central portion of the former partywall boundary and east at the Red Lake Complex. The prominent fabric at the minesite is the S2 cleavage, trending northwest southeast, axial planar of the F2 folding, plunging steeply to the south, southwest.

Several large, sill-like intrusions, ultramafic to intermediate in composition, are present in the sequence. The major mineralized zones, although hosted in basalt, are associated with a central ultramafic unit, which is a highly carbonatized and altered unit, believed to be either volcanic or plutonic in origin. This unit varies considerably from “rhyolite” breccia to talc-chlorite schist, to carbonatized and alusite-rich metasomatized rock and is characterized by intense shearing and alteration. Associated with the ultramafic is a dioritic (quartz-gabbro) unit, distinguished by the presence of blue quartz eyes.

The volcanic and sedimentary rocks and ore zones have been intruded by quartz-feldspar porphyries (QFP), metadiabase, peridotite/serpentinite and lamprophyres, post mineralisation.  The lamprophyre dykes occur in a conjugate set with a south-southeast trend dipping at 55-65º to the west-southwest, and a south trend dipping 20-40º to the west.

8.3.1       Mineralogy

The mineralogy of the mafic volcanics hosting the main zones of mineralisation is principally plagioclase, quartz, fibrous amphiboles, biotite, minor chlorite, carbonate, hornblende and talc.  The altered ultramafics are mainly peridotites which are heavily silicified and potassically altered to biotite in the vicinity of the ore zones.

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8.3.2       Alteration

Hydrothermal alteration at the Red Lake Gold Mines can be subdivided into three main phases: 1) an early alteration subdivided into a) carbonatization and pervasive biotite (potassic) alteration and (b) early silicification and aluminosilicate-bearing alteration; 2) main-stage vein phase of barren dolomite to ankerite, cockade breccias, and sheeted veinlet zones with chloritic alteration; and 3)  a mineralisation phase with quartz-sericite+/- cordierite alteration and a late episode of veinlet controlled biotite +/- tourmaline alteration.

8.3.3       Structure

The Red Lake Gold Mines mineralised system is a wedge-shaped zone above roughly 27 level which widens upwards and is constrained by bounding fault structures on the northeast and southwest flanks.  This wedge is defined by steeply south-dipping and S- to SSW- plunging litho-structural packages of ultramafic and rhyolitic bodies, enveloped mainly by metabasalts. These bodies outline dismembered folds which also plunge steeply S to SSW. Fold hinges are preserved between a series of steeply S-dipping and NW- to WNW-striking fault zones (see Figure 8.5). Separating the detached folds, these major curviplanar fault/shear zones developed in fold limb positions and are approximately axial plane parallel.  Most major fault zones tend to steepen in dip upwards, although some faults such as the West Drift Fault, are more shallowly dipping and less curviplanar.

The Campbell Fault unequivocally located on 16 Level Red Lake Complex defines the contact between a peridotitic komatiite in the footwall position and a metabasalt in the hangingwall (see Figure 8.5). The fault zone strikes NW at 110 azimuth and dips steeply at 56 degrees SW. The fault zone exhibits kinematic indicators suggestive of a reverse component of movement and significant sinistral displacement.

The Campbell Fault system is situated in the hangingwall of the South Cross-cut (SXC) at surface. The SXC strikes NW and dips 50 – 60 degrees to the SW. The SCX terminates the Campbell Fault at depth, and continues downwards subparallel to the West Drift Fault (WDF).

The WDF strikes NW and is relatively shallowly dipping (50 – 60 degrees SW). It has been interpreted in the lower levels of the mine as a hangingwall splay from the Campbell Fault.

Slightly more North-Northest (NNW) - Northwest (NW)-striking fault than the 01 or the 70 faults and generally shallower dipping (50 – 60 degrees) to the SW, The WDF has an apparent dextral displacement of approximately 60 meters, which has been interpreted at the Campbell Complex as approximately 300 meters reverse displacement based on the reconstruction of mineralised komatiite contacts.

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The WDF terminates the 70 Fault and the 01 Fault and appears to trim the high grade mineralised zone into lower grade mineralisation in the footwall.

The 70 Fault Strikes NW and dips steeply (70 degrees) to the SW. The fault on 18 – 24 level Campbell Complex where it passes between the P-zone and L-zone has an apparent dextral displacement of approximately 200 meters decreasing to approximately 60 meters in the upper levels. The 70 fault splays from or joins the hangingwall of the West Drift Fault (WDF) between 20 Level and 27 Level).

The 01 fault strikes NW and is steeply dipping to sub-vertical to the SW. An apparent sinistral displacement of 800 m is estimated for this fault and has been interpreted as equivalent to the Red Lake Fault at the Red Lake Complex. Apparent displacement is inferred from displacement of hangingwall rhyolite against footwall ultramafic and basalt. The stratigraphy in the hangingwall of the 01 fault is not as well understood as in the footwall.

Black Line Faults (BLF). Narrow veins (1 mm – 2 cm wide) filled with a dark silicate mineral (probably tourmaline) show smaller scale displacement. In two locations on 27 Level (near the shaft and 27 sub-level 2) the black line faults were intensely folded (similar to the folding of the carbonate veins). This indicates that the BLF were formed during or after an episode of folding. Other BLF zones have been observed to overprint the mineralisation.

The Kovala Fault exists in the hangingwall of the HGZ lenses, the HW5 and HWA ore structures. It appears to intersect the westerly plunge of the HGZ below 47 Level, and deep drilling suggests the ore structures experience minor attenuation prior to passing into the hangingwall.

9.0          DEPOSIT TYPES

The mineralisation at Red Lake may be classified as an Archaean greenstone-hosted quartz carbonate vein deposit (GQCV).  These typically occur in deformed basalts and ultramafic komatiite flows intruded by intermediate to felsic porphyry intrusions, and sometimes with swarms of albite and lamprophyre dykes.

The following description has been drafted with primary reference to the classifications contained within Mineral Deposits of Canada by Dubé and Gosselin (2000).

The GQCV deposits correspond to structurally controlled complex epigenetic deposits characterized by simple to complex networks of gold-bearing, laminated quartz-carbonate fault–fill veins. These veins are hosted by moderately to steeply dipping, compressional brittle-ductile shear zones and faults, with locally associated shallow-dipping extensional veins and hydrothermal breccias.  The deposits are hosted by greenschist to locally amphibolite facies

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metamorphic rocks of dominantly mafic composition and formed at intermediate depth (5–10 km).  The deformation is syn- to late deformation and is typically post-peak greenschist facies or syn-peak amphibolite facies metamorphism.  They are typically associated with iron carbonate alteration.  Gold is largely confined to the quartz-carbonate vein network but may be present in significant amounts within iron-rich sulphidized wallrock selvages or within silicified and arsenopyrite rich replacement zones.

There is a general consensus that the GQCV deposits are related to metamorphic fluids from accretionary processes and generated by prograde metamorphism and thermal re-equilibration of subducted volcano-sedimentary terranes.  The deep-seated, gold-transporting metamorphic fluid has been channelled to higher crustal levels through major crustal faults or deformation zones.  Along its pathway the fluid has dissolved various components, notably gold, from the volcano-sedimentary packages, including a gold-rich precursor.  The fluid then precipitated as vein material or wallrock replacement in second and third order structures at higher crustal levels through fluid-pressure cycling process and temperature, pH and other physico-chemical variations.

The diagnostic features of the greenstone-hosted quartz-carbonate vein type gold deposits are arrays and networks of fault and shear-zone related quartz-carbonate laminated fault-fill and extensional veins in associated carbonatized metamorphosed greenstone rocks.  The deposits are typically associated with large-scale (crustal) compressional faults.  They have a significant vertical extent (~2kms), with very limited metallic zonation.  They can coexist regionally with iron-formation hosted vein and disseminated deposits as well as with turbidite-hosted quartz-carbonate vein deposits.

Greenstone hosted quartz-carbonate vein deposits are one of the most significant sources of gold and account for more than 13% of all world gold.  Of the 14 world class deposits of this type within Canada, 6 are found within the Abitibi greenstone belt and include Timmins, Kirkland Lake and Larder Lake within Ontario.

The following table 9.1 compares the geological characteristics of the Red Lake area to other Greenstone hosted quartz-carbonate vein gold deposits:

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Table 9-1 Summary of geological charecteristics of the Red Lake Gold Mines and generic
	GQVC deposits	Red Gold Lake Mines
Host Rocks	tholeiitic basalts and ultramafic komatiite flows	tholeiitic basalt to peridotitic komatiite
Age of Mineralisation	Archean (mainly), also Proterozoic and Paleozoic	Archean
Deformation History	collisional or accretionary orogenic event	collisional or accretionary orogenic event
Alteration Style	greenschist to amphibolite facies metamorphism	The upper levels of the Campbell and western Red Lake Complexes are greenschist facies grading into lower amphibolite facies at their lower levels
Mineralisation Style	Au associated with biotite/phlogopite, amphibole, diopside, pyrite, pyrrhotite and/or arsenopyrite	sulphides (arsenopyrite, pyrite and pyrrhotite) and gold commonly accompany quartz-rich infill in re-opened veins
First-Order Structural controls	Compressional crustal scale fault zones	The movement on major faults controlled the initial alteration and orebody location and development.
Second-Order Structural Controls	Sheeted and stockwork veins associated with shear zones, faults and extensional veins	The ore veins occupy tension fractures in folded and faulted pre-Cambrian volcanic rocks. The veins generally strike to the north-west and dip steeply to the south. These zones are sub-parallel to the hosting deformation
Genetically-Associated Intrusives	Intermediate to felsic intrusives	rhyolite to dacite dykes
Average grades	5 to 15g/t Au	12 to 65g/t

The Red Lake Gold Mines area is dominated by greenschist grade rocks which have experienced intense alteration in the form of Fe-carbonatization, silicification and sericitization.  The area is also characterized by a strong component of brittle deformation in the form of intense fracturing and small-scale faulting.  Gold mineralisation has occurred late in the geological history of the area, i.e., post folding and post Fe-carbonatization.

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Fold axes that trend northwest and north to northeast indicate at least two directions of principal stress during the formation and deformation of the Red Lake belt.  The dominant northwest trending, steeply southwest dipping foliation is defined by carbonate veinlets, narrow zones of silicification, pillow elongation, sulphide streaks and a well-developed foliation in the ultramafic rocks.  Less pervasive north trending, west dipping structures defined by early carbonate veins and later gold-bearing silicification, pressure solution residue and fault structures indicate both pre- and post-foliation compression from the east.  The overall structure of the mine area can be described as a dome.

 Faults and shear zones were developed contemporaneous with folding, resulting in block movement along a mainly northwest-southeast trend.  The movement on the major faults controlled the initial alteration and orebody location and development.  Much of the movement was of a cyclical nature, constantly brecciating and resealing the structure before and during gold deposition.

10.0        MINERALISATION

In general, there are three types of mineralisation zones encountered at the Red Lake Gold Mines, namely vein type ore, disseminated sulphide ore and replacement ore. Structures at the mine exhibit three trends: conformable northwest, north-south and east-west. The conformable structures are most common and are subparallel to the foliation. The vein systems follow these structures. Complex vein arrays are those which also include the north-south and east-west components. The arrays are most common near high angle mafic-ultramafic contacts. The High Grade Zone occurs in such an environment where enhanced dilatency developed and was sustained over a long period of time. Its geometry will combine both conformable and complex vein arrays overprinted by replacement mineralisation.

The ore veins are normally structurally controlled; averaging 1.5 to 1.8 metres in width and extending over strike lengths ranging from 30 to 300 metres. Sulphide replacement zones vary from 3 to 12 metres in width and extend over a strike length of 120 to 180 metres (Figure 10.1.

Gold mineralisation zones in the Balmer Assemblage of the Red Lake Gold Mines system can be broadly subdivided into two morphological groups: Planar to curviplanar zones and plunging zones.

10.1        Planar to curviplanar zones

These have strike lengths and dip extents of a three hundred meters, but comparatively narrow widths.  These zones mainly dip moderately to steeply south and strike WNW-NNW, mainly

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subparallel to the main (S₂) foliation. Two distinct styles of planar/curviplanar zones are recognized:

Quartz-carbonate vein systems (eg ‘A’ and ‘F’ zones in Campbell Mine) comprising relatively continuous vein sets of locally variable thickness, enveloped by foliated hosts typically showing strong carbonate-biotite alteration.  Mineralized quartz-carbonate vein systems are commonly best developed in metabasaltic rocks, but also occur in dioritic, rhyolitic and ultramafic hosts.  The auriferous mineralisation is less continuous than its quartz-carbonate vein envelopes, and ranges from low to high grade, the latter consistently occurring in southwest plunging shoots.  Elevated gold grades are associated with quartz-rich infill in re-opened veins, forming lenticular zones subparallel to vein margins, as well as infilling brecciated veins and filling high angle fracture zones. Siliceous replacements of carbonate are also common and exhibit varying proportions of pyrite, arsenopyrite, stibnite and pyrrhotite. Quartz-carbonate gold systems are typically non refractory.

“Sulphidic”, replacement-dominated mineralisation zones (eg TP zones and M & M in Campbell.) occur in sulphide-, quartz- and silicate (biotite⁺/_- amphibole) + magnetite assemblages mainly over-printing carbonate -altered metabasaltic hosts and less commonly in metasediments, and typically exhibits fewer carbonate and siliceous veins than other styles.  Although described as ‘sulphide ores’ such mineralisation typically contains .5-4 percent sulphide by volume, mainly pyrrhotite and pyrite (+ arsenopyrite).  Sulphide-rich mineralisation is commonly of low to moderate grades (10-15g/t Au) and refractory in character, but may occur over substantial widths (three meters), strike lengths and depth extents.

10.2        Plunging zones

These zones have roughly equant to NW-elongate footprints up to thirty meters across and with pipe-like to linear geometries, which may extend three hundred meters down plunge.  Examples include Campbell ‘G’ zone. These comprise geometrically complex arrays of sulphidic pod-replacement-and siliceous vein mineralisation, which overprint carbonate-quartz vein networks and carbonate alteration.

Plunging mineralized zones are preferentially hosted by metabasalts, although some smaller bodies occur in ultramafics. All are spatially associated with mafic-ultramafic contacts and are localised in relatively steeply plunging hinge zones/noses of dismembered folds, where these abut major fault/shear zones. Plunging ore zones may “tail off” into planar mineralized zones, which may contain some magnetite.

Plunging ore zones may exhibit large tonnages of medium grade, low cost mineable material (eg Campbell “G” zone) and also very high grades (e.g., Red Lake Mine HGZ). Much of the ore

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is non-refractory, high grades being associated with strong silicification, arsenopyrite development and quartz veining.  These higher grade plunging, equant zones have involved more intense siliceous, arsenopyrite -rich replacement and higher degrees of dilation relative to lower grade examples and also planar mineralised zones.

10.3        Other

Gold mineralisation has also been reported in Bruce Channel metasedimentary hosts. In some examples this mineralisation is spatially associated with graphitic rocks and in others with sulphidised iron formations.

10.4        Zones

10.4.1     Red Lake Complex Zones

The High Grade Zone (HGZ) consists of quartz-carbonate veins and breccia structures and arsenopyrite replacement mineralisation within altered basalts and altered ultramafic rocks. The alteration consists of chlorite, biotite, silica, carbonitization and minor actinolite. The mineralisation is characterized by consistent distribution of both coarse and fine flecks of native gold, fine acicular arsenopyrite and pyrrhotite.  Accessory mineralisation includes chalcopyrite and sphalerite. Stibnite has been noted in only a handful of drillholes. The HGZ is comprised of several distinct lenses including the prolific Main, East-West, multiple FW zones, HW5, HWA, and HW7. Below 37Level the HW5, HWA and HW7 are the most prominent HGZ lenses. Geometries are complex, but range from North-South, to mine trend northwest-southeast to East-West orientations. The HGZ exhibits an average strike length of approximately, 60 meters, and a incredibly consistent plunge extent of approximately 1000 meters, from 30 to 49 Level.

Mining after the strike has focused primarily below 30 Level. Seventy five percent of the remaining resources above 23 Level are situated in the No. 1 shaft area and were the focus of historical mining from 1948 to the start of the strike in 1996 (figure 10.1). The majority of the mined out mineralization included quartz-carbonate vein structures. Near surface the majority of the vein structure occurred in close proximity to the Campbell Fault system including the B and Upper E Zones, the NC and the D Zones. The SC-ESC structure is laterally equivalent to the A Zone at the Campbell Complex. The E Zone, more of a replacement style disseminated sulphide mineralization with the F Zone vein in the hangingwall occurs in proximity to the former boundary with the Campbell Complex between roughly 17 and 30 Levels. Farther to the east the PLM Zone, potentially the up-plunge extension of the FW Zones and the Far East Zone remain largely untested resources in the No. 2 shaft area.

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The footwall (FW) Sulphide Zones contain disseminated gold-bearing pyrrhotite and pyrite mineralisation found in combination with replacement type ore.  The replacement refers to areas that have been permeated or flooded with goldbearing silica (± sulphide). Mineralogically, the replacement zones can be quite variable. Between 30-40 Levels the FW sulphide ore lenses are comprised of the ESC-3J, SC-ESC, and the ESC-HW.

Vein mineralisation has been dominant in the upper levels at the northwest end of the mine. With depth, vein dips become shallower and change in mineralogy from quartz-carbonate-arsenopyrite-pyrite to quartz-carbonate-pyrrhotite-arsenopyrite assemblages.

10.4.2     Campbell Complex Zones

The main mineralized zones at the Campbell Complex are the A, A1, AU, F, F2, L, NL, G, P,  DC, TP, MM and N zones. The A, F, F2, 56, DC, TP and MM zones occur along a northwest-southeast foliation, and dip to the southwest at 70 to 80 degrees. The G, L, P, NL zones are formed along the contact of the central ultramafic unit and basalt units. (figure 10.1)

The A, A1 Zones are a foliation parallel type vein system which varies from 1 feet-3 ft (0.3 to 1m) wide and consists of quartz and quartz/carbonate material with pyrite, pyrrhotite, arsenopyrite, and gold.  The zone is approximately 1500 ft (460m) long and extends from surface to the 26 level (3850 ft) and is located on the hanging wall (HW) side of the central ultramafic rock unit, south of the shaft.  The A1 Zone is a splay of the A Zone on the HW side with similar mineralisation and is 750 ft (230m) long and extends from the 13 - 21 level.  The A Zone corresponds to the Red Lake Complex ESC Zone.

The AU Zone is a foliation oblique, near vertical, pipe-like, vein and replacement style mineralised system plunging to the south-west which varies from 1 ft-10 ft (0.3m – 3m) wide. It consists of quartz lenses and pipes containing pyrite, pyrrhotite, arsenopyrite, gold, sphalerite, stibnite and fuchsite.  This type of zone varies in length from 10 ft-250 ft (3m – 75m).  The AU Zone is located south of the No. 1 shaft.

The B Zone is a foliation parallel type vein system averaging 5.0 ft (1.5m) wide that extends up to 900 ft (275m) in length on the central levels between the L Zone and the G Zone and vertically from the surface to the 24 level.  The B Zone consists of quartz and quartz/carbonate material with pyrite, pyrrhotite, sphalerite, cinnabar, fine arsenopyrite and gold.  The B Zone is located on the footwall (FW) of the central ultramafic between the L and G Zones and north of the No. 1 shaft.

The DC Zone is comprised of multiple stacked vein structures and localized high grade 10-20ft (3-6m) podiform structures. The structure is essentially quartz-carbonate veins with arsenopyrite

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replacement mineralisation. The alteration minerals consist of chlorite, biotite, silica, carbonitization and minor actinolite.  The DCE Zone potential strike extension corresponds to the HGZW of the Red Lake Complex along the former partywall boundary.

The F Zone is a foliation parallel type vein system which varies from 1 to 3 ft (0.3 – 1m) wide and is approximately 1000 ft (300m) in length and extends from surface to the 19 level.  The zone consists mainly of quartz with lesser amounts of quartz/carbonate and is mineralized with pyrite, pyrrhotite, sphalerite, stibnite, fuchsite, and gold.  It is located west of both shafts and south of the G Zone.  It is bounded by the rhyolite to the west and intersects the central ultramafic unit in the east.

The F2 Zone is a foliation parallel type vein system which varies from 1 to 3 ft (0.3 – 1m) wide. It is approximately 600 ft (180m) in length and extends from the 2 level to below the 27 level.  The zone consists mainly of quartz and quartz carbonate and is mineralized with pyrite, pyrrhotite, sphalerite, stibnite, fuchsite, and gold.  It is located west of the No. 1 Shaft and south of the G Zone near the HW of the rhyolite and west of the A Zone along strike.  The zone is cut off by major faults on the west end and intersects the central ultramafic unit on the east end.

The G Zone is a foliation oblique type vein system approximately 500 ft (150 m) long and varies from 10 to 70 ft (3m – 21m) wide.  This zone extends from surface to 22 level and although the structure continues at depth the vein is uneconomic.  The zone consists of quartz and quartz/carbonate with silicified replacement sections overprinting the carbonates and is mineralized with heavy pyrite and pyrrhotite, arsenopyrite, sphalerite, stibnite, fuchsite, and gold.  The G Zone is located north of both shafts at the contact of the basalt (andesite) and a nose in the north (ultramafic) unit.

The H Zone is a series of foliation parallel veins which varies from 1 to 3 ft (0.3m – 1m) wide and is approximately 1400 ft (430m) in length.  This zone extends from the 3 level to the 10 level and is open to depth.  The zone consists mainly of quartz with lesser amounts of quartz carbonate and is mineralized with pyrite, pyrrhotite, arsenopyrite, and gold.  The H Zone is located west of the G Zone and north of the Reid Shaft.

The L Zone is a complex series of foliation parallel to foliation oblique veins situated along a fold in the central komatiite-mafic contact, varying in length from 50 to 150 ft (15 – 45m) and striking northwest. The main L Zone vein is approximately 300 ft (90m) in length and varies in width from 1 to 120 ft (0.3 to 35m) and extends from surface to the 22 level. Additional veins vary in length from 50-150 ft (15-45 m) and upto 10ft (3 m) wide The zone consists of quartz and quartz/carbonate with silicified replacement sections overprinting the carbonates and is mineralized with pyrite, pyrrhotite, arsenopyrite, sphalerite, stibnite, fuchsite, and gold. This zone is located east of the No. 1 Shaft and most resembles a complex stringer/stockwork zone.  The

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North L Zone is a foliation parallel vein structure similar to the A zone, 1-3 ft (0.3-1 m) in width and 200-300 ft (60-90 m) in strike length, extending from surface to 22 Level, where the gold grade becomes uneconomic.

The MM Zone is a sulphidic replacement-dominated mineralisation zone with a sulphide, quartz and silicate (biotite⁺/_- amphibole) + magnetite assemblage mainly over-printing carbonate -altered metabasaltic hosts and less commonly in metasediments.  It typically exhibits fewer carbonate and siliceous veins than other styles.

The N Zone is a foliation parallel vein in the basalt and foliation oblique vein in the ultramafic rock and varies from 5 to 20 ft (1.5 - 6m) wide and 100 to 250 ft (30m to 75m) long.  The zone is located west of the shaft between the L & P Zones and extends from the 14 level to the 27 level.  This zone is similar to the AU with half the ore in andesite and the other half in altered rock.

The O Zone is a foliation parallel type vein system which varies from 1 to-5 ft (0.3 – 1.5m) wide and approximately 350 ft (105m) long.  The zone is located north of the shaft and the G zone and is developed on the 11 and 12 levels.  The zone consists mainly of quartz/carbonate and is mineralized with pyrite, pyrrhotite, arsenopyrite, and gold.

The P Zone is a combination of foliation parallel veins and foliation oblique type veins approximately 450 ft (140m) long which vary in width from 1 to 65 ft (0.3 – 20m), extending from the 18 level to the 27 level.  The zone has three main veins radiating west from the basalt-ultramafic contact.  It is situated west of the shaft, south of the G Zone and west of a nose in the central ultramafic rock unit.  The zone consists of quartz and quartz/carbonate with silicified replacement sections overprinting the carbonates and is mineralized with pyrite, pyrrhotite, arsenopyrite, sphalerite, stibnite, fuchsite, and gold.

The TP Zone and PCB Zone are characterized by intense silicification, arsenopyrite, pyrite and/or pyrrhotite, and base metal sulphide mineralisation ranging from <5% total sulphides to semi-massive sulphides. These replacement type mineralised zones extend a thousand feet into the footwall of the 56, 56-1 vein structures, and the DC Zone to depth.

The 56 and 56-1 zones are both horizontally and vertically extensive, striking for approximately 2000 feet (600 m) and extending vertically for 1800’ (550 m). The 56 zone is a basalt hosted quartz-carbonate vein and is typically between 2 to 4 feet (0.6-1.3 m) wide. Auriferous mineralization may be present where strong silicification and sulphide mineralization (pyrite and pyrrhotite) occur with the vein

The 56-1 auriferous structure (hangingwall to the 56 vein) is hosted in both basalt and komatiite lithologies. Within the basalt the structure may consist of one ore more narrower mineralized

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veins, commonly 0.5 to 1.0 feet (0.15-0.3 m) wide. The structures are generally auriferous with both grade and width decreasing away from the basalt-komatiite contact. Ore grade mineralization is typically found proximal to the contact or within the komatiite unit. Within the komatiite, gold mineralization is found where strong carbonate and biotite alteration, and abundant veining is present.  Ore mineralization is associated with strong silica replacement and suphide mineralization (predominantly aresenopyrite).  Visible gold is common, with ore width commonly from 2’ to 4’ (0.6-1.3 m) wide, although widths of 6’ to 8’ (1.8- 2.3 m) aren’t uncommon.

11.0        EXPLORATION

Integration of the Campbell and Red Lake Complexes has provided the geology team with a unique challenge combining the strengths of grade control and reconciliation of the Red Lake Complex and the long term planning and resource block modeling process at the Campbell Complex. Of the many tangible benefits of the merge, the resultant exploration portfolio has to rank very high on the list of synergies. One of the key early developments was the bolting together of the two 3-D geologic block models on the GOCAD platform at the virtual reality Lab. Investment in Technology and academic research has resulted in a strong geologic model that will pay dividends in the future.

As a service provider for production, the fundamental objective in exploration geology is to strike the right balance of exploration, delineation and definition drilling to facilitate the replacement of mined out reserves, fine-tune ore outlines on development attack drifts for cut and fill stopes, sublevel developments for longhole mining blocks, and discover new mining horizons that can provide growth opportunities to Life of Mine production profiles. These objectives require close alignment with the Mineral Resource and Mineral Reserve definitions and classifications of measured, indicated and inferred (Figure 11.1). A steady stream of upgrading from inferred to indicated class is required, since the engineering group requires at least indicated class confidence level to build mine plans converting resources to reserve category. This confidence level is a function of drilling density, and depending on the nature of the mineralisation requires 25-50 foot (7.5- 15 m) spaced drill centers to warrant indicated drill status, assuming the “QP” is comfortable with the geologic continuity of the ore structure.  In fact, measured and indicated resource classes require some level of development drifting on structure providing muck and chip sampling to effectively qualify as a reasonable prospect for economic extraction. Grade control sampling provides the geologic reconciled grade to the grade estimation process highlighting the importance of data flow between exploration and production geology in the pursuit of quality MRMR.

Despite a prolific mining history since 1948-49 at the RLGM, minimal exploration activity and development were focused in proximity to the former boundary, which appears to transect the

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centroid of the orebody (Figure 11.2).  Recently developed inter-mine connections (i.e., 36/34 Level transfer) provide operational flexibility and staging platforms for exploration of these partywall targets. Deep exploration depends heavily upon critical development away from the orebody for quality evaluation of highly sought after high grade vein structures to depth.

The current high priority exploration targets include the following: (i) the HGZ and associated footwall sulphides, (ii) the Deep Campbell zone, (iii) ‘Party Wall’ (or internal property boundary) opportunities, (iv) the Upper Red Lake Sulphides, and (v) surface, bulk mining opportunities (Figure 11.3)

The exploration targets are characterized by relatively smaller (i.e. by volume) high grade vein structures and laterally extensive footwall, lower grade, sulphide ore structures.

11.1        HGZ and FW Sulphides

The HGZ has been the focus of mining and exploration since the labour lockout at Red Lake Complex between 1996-2000. Deep drilling collared on 34 Level footwall position and subsequently the 37 Level hangingwall position employed devico controlled drilling techniques to steer holes onto the remarkably consistent westerly plunge of the HGZ between 30 Level and 49 Level.

Grade estimation is achieved using bounded 2-D polygonal drapes on a 3-D solid. The reserve has been drawn down to 47 Level and the resource down to 49 Level (Figure 11.4). Although the HGZ is comprised of multiple discrete lenses, below 37L the bulk of the HGZ is comprised of the HW5, HWA and HW7 lenses. For resource classification all intercepts are diluted to a minimum horizontal width of 4 feet and capped at mean plus four standard deviations.

In 2004 with the merger of Wheaton River Minerals, deep exploration drilling targeting the HGZ was suspended until development platforms facilitating more cost effective drilling were in place. In the fall 2005, a drilling campaign from 37 Level Red Lake Mine targeted the down plunge projection of the HGZ onto former Campbell Mine property was initiated by Placer Dome (CLA) Limited.  Economic intersections at the 57 Level elevation were returned in two of three holes completed roughly where the plunge line projected. Although geologists at the Red Lake Mine were undecided on whether or not the intersections were unequivocally HW5 or HW7, the existence of ore bearing structures to depth is extremely promising.

Approximately 500 ft (150m) into the footwall of the HGZ, the laterally extensive orebodies referred to collectively as the FW Sulphide Zones including the SC-ESC, ESC-3J and the ESCHW lenses are hosted in a high strain corridor bounded by the Dickenson Fault (footwall) and Red Lake Fault (hangingwall).  Although the focus of mining and exploration immediately

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after the strike was concentrated on the HGZ, the attraction of the FW sulphide zones for bulk volume longhole mining methods did not go unrecognized. In an effort to reduce the dependency on the HGZ to depth, the concept of blending the FW sulphides with the HGZ was developed and a new shaft was designed with substantially increased hoisting capacity. The combined asset has taken this strategy to the next step, establishing an inter-mine connection between 34-36 Levels at the bottom of the Reid shaft to fully utilize the spare hoisting capacity and spare milling capacity at the Campbell complex in the near term prior to the completion of the No. 3 Shaft at the Red Lake Complex, targeted for January 2008.

In 2006 development into the FW Sulphides at the 34-2 sub elevation will open up the ore and provide longhole infrastructure for mining over the next couple of years (Figure.11.5). Approximately 1100 ft (335m) sublevels will refine the model shapes in preparation for longhole mining. It is anticipated that the blocks will prove substantially more continuous if the cutoff grade is dropped to 0.15-0.20 opt. Delineation drilling has focused in 2006 on increasing the drill density to 50 ft (12 - 15m) centers in the 30-40L block, while expanding the resource to depth below 40 Level. Extensive remodelling at year-end will help refine the mining strategy in 2007 and beyond.

Mining above 30L prior to the start of the strike in 1996 had reached pillar recovery stage in some relatively high grade zones ie. 0.5-1.0 opt.  Historical resources and reserves were largely ignored after the strike as mining focused on the HGZ. 75% of the remaining resource/ reserve exists in the SC, and other zones above 23 Level in the No. 1 Shaft area. Much of the longitudinal resource/ reserve data including an extensive chip dataset exist only on mylars and are currently being digitized and modeled into electronic format.

One of the advantages of the merger is that mining in the upper levels of the Campbell Complex is still very active with infrastructure relatively proximal to No. 1 Shaft ore structures at the Red Lake Complex. Drilling is currently testing the SC hangingwall on the 9 and 15 Levels Red from a drift heading in the hangingwall at the Campbell Complex situated within 500 ft (150m).

11.2        Deep Campbell

In 2006 a hangingwall exploration platform, 4199 Ramp was driven off the bottom elevation of the Reid Shaft ramp to establish a quality collar position to drill the Deep Campbell below 39 Level (Figure 11.2). Previously, inefficient 4000 ft (1220m) plus holes were collared from a footwall location near the shaft providing poor coverage of the Deep Campbell potential. The 1200ft (365m) ‘T’ drift off the end of the ramp will provide coverage on a block of ground 2000 ft (600m) in strike, 1500 ft (450m) vertical and 1000 ft (300m) thick of high exploration potential down-plunge of the DC vein structures and the overlapping replacement style PCB/TP Zones in

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the footwall. Currently a first pass program is underway with three rigs drilling on a 200 x 200 ft (60m) grid (Figure 11.5).

Mining at the lowermost mining horizon in the DC Zone has encountered multiple stacked vein structures and localized high grade “jewellery boxes”, i.e. 10-20ft (3-6m) podiform structures. Replacement style mineralisation in the footwall similar in a general sense to the FW sulphides and the HGZ comprise the TP and PCB Zones. These economically challenged structures are characterized by intense silicification, arsenopyrite, pyrite and/or pyrrhotite, and base metal sulphide mineralisation ranging from <5% total sulphides to semi-massive sulphides. These large mineralized structures associated with high deformation corridors are centered on prominent fault structures bounding the majority of know ore mineralisation at both operations. This style of mineralisation is laterally more extensive than the vein style DC Zone and occurs over an area 2000 ft (610m) in strike and 1000 ft (305m) in thickness in the footwall to the DC Zone. It is critical for the future at Campbell Complex, that the blend of the two styles of mineralisation provide an economically viable mill feed.

11.3        Partywall Opportunities

Fifty-eight years of mining history with a boundary partywall between the Campbell mine and the Red Lake mine created a substantial data void in the mid-section of the combined asset, especially below the 20 Level elevation. Infrastructure proximal to the boundary is scarce resulting in largely untested extensions of known orebodies trending towards the former boundary from both directions.

Two exciting prospects along the boundary are the E&F Zones between 20-30 Levels and the HGZW target between primarily 36-39 Levels at the convergence/ extension of the prolific 56 and DC East veins from the Campbell Complex onto the Red Lake Complex (Figure 11.6). Inter-mine connections initiated post merger provide operational synergies and exploration platforms for both of these exciting prospects.

On 24 level Campbell Complex, a drive across the former boundary is underway to cross-cut the E&F structures before year-end. The 34-36 connection between the two operations, largely complete at press-time, will provide a drilling platform for the HGZW target. The 4199 Ramp exploration platform has been extended an additional 500 feet (150 m) to the southwest in order to gain a quality platform on the down plunge extension of the HGZW below 39 Level.

11.4        Surface Exploration Targets/ Concepts

In 2005, Campbell mine site exploration over the first 1000 ft (300m) vertical was revisited because of the apparent paucity of surface drilling i.e. > 300 ft (90m) centers, extensive

- 35 - November 17, 2006

waterbody coverage that would have prevented surface prospecting, and the potential to quickly optimize the Campbell Complex mill. Conceptually, the preferred mining target was an open pit or ramp mining scenario that would facilitate the “fill the mill” mandate. A surface trenching and sampling program uncovered two potential micro-pit targets that warranted further drilling, sampling and modeling; the H Zone and G Zone North.

In the summer of 2006 the H Zone micro-pit was mined using a contractor and a pit design supplied by engineering personnel of the Porcupine Joint Venture (“PJV”), operated by Goldcorp.

A much larger pit internal scoping study was initiated in 2006 to investigate the potential of a mine plan encompassing all of the currently known crown pillar ore structures in a much larger bulk volume scenario, with expectations of a low grade halo, remnant pillars, and unknown parallel and transverse ore structures captured by the pit outline. A crude model was built upon 0.05 opt iso-shells over the first twelve levels in Datamine and linked to form a solid block model. Pit optimization run by PJV engineers returned favourable numbers on this crude model. An understanding of the background grade estimation is currently being assessed for outlined litho-structural domains by drilling and whole core sampling shallow drill traverses.

11.5        District exploration

Goldcorp is actively exploring a large land position for stand-alone gold deposits in the Red Lake District.  Goldcorp holds in excess of 64,000 ha of prospective ground in the Red Lake District, of which 42,000 ha is under 100% Goldcorp control and a further 22,200 ha under Joint Venture and Option Agreements.  The Red Lake Regional and Red Lake Gold Mines’ Exploration teams work in close collaboration to develop exploration concepts and strategies in order to identify priority exploration targets.

11.6        Additional Studies

Goldcorp has committed to a number of academic research initiatives with the Ontario Geologic Survey, the Australian Geological Survey, and the University of British Columbia MDRU (Mineral Deposit Research Unit). Sponsorship of geologic, thesis research projects at the Red Lake Gold Mines (i.e., B.Sc. to Ph.D candidates) and an embedded researcher are just some of the examples of Goldcorp’s commitment to advancing the understanding of the geologic model facilitating a higher quality exploration effort.

12.0        DRILLING

Diamond drilling has been carried out at the Red Lake Gold Mines since the beginning of their operations. Over the years various lithological descriptions have been developed for each of the

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complexes and these have continued to develop and change.  Since the merger of the two operations, Goldcorp has developed a new lithological coding system that incorporates both the old systems.  Table 12.1 gives the breakdown of how each Complex relates to the other through the new code.

Table 12.1     Reclassification of Lithological codes

GOLDCORP	NEW	DESCRIPTION	CAMPBELL
MINE	CODE		MINE
		MAFIC VOLCANICS
01	V2A	BASALT	BA
04	V2B	BLEACHED BASALT
		INTERMEDIATE TO FELSIC VOLCANICS	-
11	V3T	TUFF
12	V3T2	QUARTZ CRYSTAL TUFF
13	V3U	INTERMEDIATE VOLCANIC FLOW
05	V4A	RHYOLITE	RH
	V4C	DACITE PORPHYRY	DA
	V4U	FELSITE	FE
		MAFIC INTRUSIVES
	I3D	DIORITE	DI
08	I3D2	CAMPBELL DIORITE	CD
30	I2A	GABBRO	GB
		ULTRAMAFIC ROCKS
36	V1B	ALTERED ULTRAMAFIC (BK*)/SERPENTINITE	SE
38	V1U	ULTRAMAFIC (UNSPECIFIED)/Komatiite	KO
39	V1A	GRANULAR ULTRAMAFIC ± CARB EYES (PK*)
		CLASTIC SEDIMENTARY ROCKS
40	S1A	(GARNETIFEROUS) ARGILLITE	GA
41	S1A	ARGILLITE	AR
41	S1B	CHERTY ARGILLITE	AC
	S1F	QUARTZITE	QT
	S2	SILTSTONE	SI
	S2A	GREYWACKE SILTSTONE	GS
42	S3E	GREYWACKE	GR
43	S3F	QUARTZITE
44	S4	CONGLOMERATE	CO
		CHEMICAL SEDIMENTARY ROCKS	-
60	C1	CHERT	CH
61	C2	IRON FORMATION	IF
62	C5	MARBLE
		DYKES	-
10	I2C	ALTERED (MAFIC) DYKE	AM
	I1E	LAMPROPHYRE DYKE	LD
20	I1E1	HORNBLENDE LAMPROPHYRE
23	I1E2	MELANOCRATIC LAMPROPHYRE
	I1U	ULTRAMAFIC INTRUSION	UI

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GOLDCORP	NEW	DESCRIPTION	CAMPBELL
MINE	CODE		MINE
28	I4G	ALTERED FELSIC DYKE
29	I4P	QFP, FP DYKE
	I4P4	DACITE PORPHYRY	DP
	I4P1	BLUE QUARTZ PORPHYRY	BQ
	I4P3	FELDSPAR PORPHYRY	FP
	I4P2	QUARTZ FELDSPAR PORPHYRY	QF
		ORE / VEIN / MINERALISATION TYPES
	M	MINERALIZED ZONE	9M
	M1	VEIN	9F, 9L, 9S, 9V
70, 7071, 7072,	M1A	CARBONATE-QUARTZ VEIN / BRECCIA
7076, 7077		(>50%)
71, 7170, 71707,
7172, 71727,	M1C	QUARTZ VEIN
7176, 7177
	M2B	REPLACED ZONE	9N
	M3	REPLACED HOST	9R
72, 7270, 7271,	M3A	REPLACEMENT (aspy + silica)
7276, 7277
76, 7670, 7671,	M3B	MINERALIZED (sulphides)
7672, 7677
77, 7770, 7771,	M3C	MAGNETITE REPLACEMENT / vein
7772,7776
79	M4	GOLD within SULPHIDES
		MISCELLANEOUS
LC		LOST CORE	MC
CS		CASING	CA
CL		CLAY
GV		GRAVEL
OR		ORGANICS
SD		SAND
ST		SILTSTONE
		OVERBURDEN	OB
		OTHER	OT
		STRUCTURE
50	T1A1	BLACK LINE FAULT
51	T1B	GOUGE FAULT
52	T1C	CARBONATE BX FAULT
54	T2	SHEAR ZONE
	T3A	TALC CHLORITE SCHIST	TC
55	T9A	JOINT (no infill)
57	T9B	JOINT (calcite)
59	T9C	JOINT (gas-staining)

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12.1        Red Lake Complex Diamond Drilling

The Mineral Resources and Reserves at and above the 30 Level are based on over 21,000 surface and underground drillholes.  Currently, the below 30 Level database for the High Grade Zone and the Sulphide Zones contains about 5,300 diamond drillholes totalling more than 2.5 million feet of drilling, predominantly AQTK wireline, roughly 1.2 “ (30.4 mm) in diameter.

Definition drilling below the 30 Level has been concentrated within a block of ground 4,000 feet (1220 m) long and 4,000 feet (1220 m) high, which extends to about 8,000 feet (2440 m) below surface.  The principal accesses for drilling the High Grade Zone are the 30 Level, at the base of the main mine workings, 34 Level and 37 Level, as well as the 32, 33, 35, 36 and 38 sub-levels.  Bazooka-type drills are used for detailed definition drilling from within the stopes when access is available.

The general objective of the exploration/definition drilling program has been to outline the mineralisation at 100-foot (30 m) centres and to detail and delineate the Mineral Reserves at 25 to 50-foot (7.5- 15 m) centres for mine planning.  Much of the known mineralisation has now been defined at 50-foot (15 m) intervals, but intercept spacing is variable due to the irregular location of drill sites and the complex distribution of the mineralized zones.  Goldcorp attempts to complete this in-fill using shorter drillholes from closer range.

Deep drilling on the plunge of the HW5 and HWA zones is being done from the 34 and 37 Levels, and intercepts have now been obtained below 50 Level with holes in excess of 4500 feet (1370 m) in length.  Deep exploration drilling is normally conducted with Boyles 20 drill rigs using NQ wireline size core, 1.8 “ (48 mm) in diameter. In order to reach the target area, hole deviation is accurately surveyed and the direction of the holes is controlled using Devico deflection drilling tools, which provide continuous core throughout.  Drilling and survey guidance is performed by specially trained drill crews.

Deep drilling is hampered by soft altered ultramafic (komatiite)  rocks and exploration of the downward extension of the High Grade Zone is limited to only the upper members of the Hanging Wall group.  The downward projection of the upper portion of the High Grade Zone below the 37 Level continues to be delineated by systematic drilling.

12.1.1     Drill Core Logging Procedures

Exploration drill core is transported to the core facility in sealed boxes.  Upon arrival it is marked up by a geologist and then geologically logged while wet.  All drill core is logged using computer codes for the various rock types, mineralisation, alteration characteristics and structural/geotechnical data.  The shear structures containing the various mineralized zones are

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logged in detail to establish the zone width and sampling interval.  Sections of core with high-grade visible gold mineralisation are transferred into safe locked storage to ensure that no core is lost or replaced, prior to shipping for assay.

12.1.2     Mine Grid and Survey Control

The collars of all drillholes are surveyed by transit for location, bearing and dip and tied into the mine grid.  Downhole surveys are currently (since 1995) being conducted in a systematic manner with the Gyroscopic survey instrument (unaffected by magnetics) for holes steeper than 70°, and the Reflex Maxibor survey instrument for holes with flatter dips.  Project specifications require downhole surveys at 100-ft (30m) intervals or less.  In the earlier stages of the project, Sperry Sun, Icefield multishot, Light-Log and Tropari instruments were also used, but the Gyro and Maxibor units are now favoured.

There are a few drillholes where there is doubt about the intercept location However, statistical tests of the drill results indicate that any errors in Mineral Reserves due to misplaced holes would be insignificant.  Mining to-date has not encountered any problems with mislocated drill intercepts and ore outlines conform well with the polygon outlines.  Goldcorp continues to resurvey holes that appear to have location or downhole problems; however, the deviation in the drillholes is generally small and predictable.

The original surface and underground survey control of the Red Lake Complex is done set up on a local grid co-ordinate system.  The conversion from surface UTM to mine grid is as follows:

The mine grid is based on UTM NAD 27.

•18553967.05 ft + Mine Northing = UTM ft Coordinates

•1464354.40 + Mine Easting = UTM ft Coordinates

•Mine Elev - 8762.09 = Elev Above Sea Level in Ft

The difference from true north and magnetic north is Declination is 0 Deg, 38 min East

The GPS equipment used is Topcon’s Hyper, dual constellation, real time kinematic system (RTK), which is a DGPS (differentially corrected global positioning system). This system is a two piece GPS ensemble capable of measuring within millimetres the distance to the orbiting GPS satellites and is composed of a base unit (stationary) and a rover unit (mobile).

The base unit is established over known survey control stations and transmits satellite corrections to a rover unit. The rover GPS antenna is located on the top of the surveyor pole.

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By receiving the base GPS correctional signal, the rover unit is capable of 10-millimeter and 15-millimeter relative accuracy from the base in both horizontal and vertical respectfully.

In order to orient the GPS system to the local grid it needs to be calibrated and oriented in the same direction and orientation as its local grid co-ordinate points.  This is achieved by setting up the GPS unit over two or more known control points.  The locations of the control points used can be seen in Table 12.2.

Table 12.2     Location of Mine Surface Survey Points

Point ID	Northing-Mine Gr	Easting-Mine Gri	Elevation-Mine Gri	Mine Site	Description
st-f	5097.841	5829.841	1279.838	Campbell	Near B-Zone Fan
st-k	9056.085	4230.391	1215.4	Campbell	HG-Young
st-t	6086.534	6298.871	1265.4	Campbell	NE Property area
st-v	6781.251	5753.708	1244.94	Campbell	Near Powder Mag

12.2        Campbell Complex Diamond Drilling

The drill hole database consists of a total of 23,761 diamond drill holes, drilled from underground and surface The channel sample database consists of a total of 81,074 channel cuts.

All underground definition and delineation drilling is AQTK wireline roughly 1.2 inches (30.4 mm) in diameter diamond core. Exploration drilling is usually BQ or NQ size core, 1.4 to 1.8 “ (36.5-47.8 mm) in diameter. Drill core sample security is maintained throughout the year with geological supervision of transport of the core from the underground/surface drill site, through to the logging facility and to the in-house assay laboratory.

Underground delineation drill spacing is based upon an approximate 10 ft – 50 ft (3m – 15m) interval spacing with more detailed drilling in select areas.

All drill hole collars are surveyed by mine surveyors, and down hole surveys are taken either with a Maxibore survey instrument at 5 foot intervals or an acid test/Tropari every 100 ft (30m). Most of the drill holes greater than 400 ft (120m) are surveyed using the Maxibore method.  All stopes will have channel sampling taken every 8-15 ft (2.5 – 4.5m) to fill in the gaps from diamond drill intercepts. More recently Reflex and Ranger electronic compass single-shot surveys tests are conducted every 100-150 ft (30-50 m)  down the hole, especially for drill holes exceeding 500 ft. (150 m) in length.

12.2.1     Drill Core Logging Procedures

Half of the core is kept for the majority of exploration and all grassroots exploration drill programs.

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All holes are logged for lithology, alteration, structure and veining/mineralisation and RQD where applicable.

12.2.2     Mine Grid and Survey Control

The original surface and underground survey control of the Campbell Complex is set up on a local grid co-ordinate system.  The conversion from surface UTM to mine grid is as follows:

The mine grid is based on UTM NAD 27.

•18553967.05 ft + Mine Northing = UTM ft Coordinates

•1464354.40 + Mine Easting = UTM ft Coordinates

•Mine Elev - 8762.09 = Elev Above Sea Level in Ft

The difference from True north and magnetic north is Declination is 0 Deg, 38 min East.

13.0        SAMPLING METHOD AND APPROACH

13.1        Red Lake Gold Mines (Red Lake and Campbell Complexes)

Rigorous sampling is done during production for grade control and correlation with Mineral Reserves estimates and production.

At the Red Lake Complex, muck and chip sampling is performed on a blast by blast basis by the production geology team, while muck sampling at the Campbell Complex is done by the miner during the mucking process.  Muck samples are used to provide a general guide and back-up information for day to day operation, while testholes are required to ascertain that no mineralisation is missed in the walls of the stope.

The reconciled grade on a stope by stope basis is variable between reserves and actual production.  The reconciliation table below demonstrates variability between muck, chip and drill results for the Red Lake Complex.

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Table 13.1     Comparison of 2005 Mining Reconciliation vs Drilling, Chip and Muck Samples

(to December 31, 2005)

Level	Actual		Drilling		Chips		Mucks
	Tons	Grade	Uncut	4 SD	Uncut	4 SD	Uncut
31	4,943	1.32	0.59	0.59	1.29	1.18	1.36
32	8,922	1.22	2.36	2.18	1.51	1.19	1.82
33	60,325	2.02	1.78	1.69	3.15	2.14	2.3
34	67,762	2.08	2.97	2.67	3.02	2.55	2.72
35	36,214	2.07	2.51	2.27	3.91	2.79	2.66
36	39,346	2.92	4.28	3.94	5.8	4.11	4.01
37	648	2.49	2.57	2.41	6.21	2.84	3.75
38	19,652	3.52	3.16	3.16	5.54	4.34	4.77
39	28,430	3.38	2.76	2.74	4.86	4.29	5.34
Total	266,242	2.39	2.73	2.55	3.88	2.96	3.05

13.2        Chip Sampling

All chip samples are taken either by a geologist or an experienced sampler.  A weighted average grade is determined for each blast based on the assay results of those samples influencing the grade of the volume blasted.    These samples are most often collected at the mid-lift elevation.  Occasionally, wall samples are also used to determine grade when the geometry of the vein dictated this usage.  The volume used to calculate the blast grade is the estimated volume preceding the face.

Although sampling guidelines are such that geologic boundaries are to be respected, the minimum sampling chip recommended is 0.5 ft (0.15 m). Additionally, where possible, 2.0 ft (0.6 m) channel chips are preferable, in an effort to duplicate the optimized drill sample interval of 2.0 ft (0.6 m). Samples along the chip sample string bracketing the mineralized structures are highly sought after to assist in the modeling of mineralized structures. Computerized modeling is facilitated by snapping to the grading selvage in contact with waste when the geologist is wireframing a 3-dimensional solid interpretation of the of the ore lens.

13.3        Muck Sampling

Muck samples are taken extensively during mining, and are collected from the majority of the ore blasts during silling and subsequent mining.  On average, at both complexes one muck sample is taken for every 20 tons of ore. At the Campbell Complex muck samples are used for reconciliation whereas at the Red Lake Complex chip samples are the predominant assay type used in reconciliation.

The chip sample and muck sample data are combined daily into the ORE REPORTER database, which is used as a monitoring centre for ongoing progress.

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13.4        Testhole Sampling

Testhole sampling is used at the mine as a grade control tool only.  Generally, testholes are 8 ft (2.4m) long and three samples are collected from each.  Testhole results are used only to identify economic mineralisation in the walls of drifts and stopes and are not used to estimate grade.  This information may result in further slashing, as required, to recover additionally defined mineralisation.

14.0        SAMPLE PREPARATION, ANALYSES AND SECURITY

14.1        Red Lake Complex

Up to July 1998, all identified mineralized structures were sampled by taking half core that was split using a diamond saw.  Geologists mark the core split using a lumber crayon optimizing the mark on the core to bisect the ellipse of the suspected mineralised structure equally. The remaining half core was saved for future reference, part of which was used for metallurgical testing.  All exploration drill core is still split at the core facility and half of it is stored in the core library on site.  However, since production began in 2000, all detailed in-fill (definition) drilling holes are sent as whole core for assaying.

Up to 1999, sample lengths were typically in the two to three foot (0.6 -1.0 m) range, and usually shorter in the higher grade sections.  Low-grade rock and waste were typically sampled over two to three ft lengths, averaging 2.20 ft (0.67m), while very High Grade sections were sampled over 0.5 to 2.0 ft (15cm to 60cms) intervals with an average of 1.60 ft (0.5m).  From 1999 onwards, sample lengths were standardized to 2 ft (65 cms) intervals, except where significant geological differences are present.

Exploration drill core is commonly assayed at outside laboratories; namely X-RAL Analytical Services (“X-RAL”), Rouyn-Noranda, TSL Saskatoon, ALS Chemex (“Chemex”), Mississauga and more recently at SGS Mineral Services (“SGS”) in Red Lake. Common standards are submitted to all laboratories.

Production chip samples typically weigh about 1 kg and core samples are 2 ft (60cms) of BQ size. All underground samples are sealed in tamper-proof bags which are transported in locked containers to surface, where check samples are added and the containers are locked again for delivery to the assay laboratory.

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14.1.1     Analysis (RLC)

Standard fire assay procedures with a gravimetric or AA finish (depending on the anticipated grade of the sample) are carried out on all assays of production samples.  All production assays are done at the X-RAL laboratory in Red Lake, while exploration samples are often done by outside laboratories.  Prior to late 1999, samples were analyzed at Chemex and more recently at SGS (after the take over of X-RAL).

Metallic screen assay methods are done at X-RAL in Red Lake for samples with visible gold or for samples where fire assay results reported more than 20 oz Au/ton.  However, there appears to be no significant difference in results between metallic screen and regular assay methods.

Exploration assay checks (duplicates and re-runs) are conducted regularly on the pulps and rejects and at the discretion of the geologist submitting the samples where assays did not match physical observations of mineralisation.  Further checks are done on rejects where the pulp assays are in doubt.  When very High Grade assays are expected, the mine staff clearly labels these samples in order to alert the lab personnel to take extra care to clean the sample prep crushers and pulverizers prior to running the next sample.

Where samples have been re-assayed, the accepted grade is the average of the two pulps, which is then averaged with the reject sample.  Check assays at outside laboratories during exploration indicated that all assays for the project were of good quality and without bias.

Samples prior to 1995 were fire assayed by Goldcorp’s onsite lab and make up the majority of the data included in the Sulphide Zones.  Late in 1999, the primary assaying contract for drill core was awarded to X-RAL in Rouyn-Noranda.  All 850 pulps from previous check assays were sent to X-RAL for comparative analysis. The results from the X-RAL checks compared well with previous checks and no discontinuity in quality of results was noted.

At the start of production in July 2000, X-RAL, now SGS, established a gold assaying facility in Red Lake that handles definition drill core and production assaying.  Exploration drill core is also done at Red Lake or sent out either to SGS in Rouyn-Noranda, TSL in Saskatoon or ALS Chemex when the lab in Red Lake cannot handle the overflow.  In 2005, samples were sent to three labs; SGS in Red Lake, SGS in Lakefield and TSL in Saskatoon, with a few RUSH production sample shipments sent to the neighboring Campbell Mine.

14.1.2     QA/QC Procedures (RLC)

Daily results for the standards and blanks are reported to the mine staff to ensure that potential problems are identified immediately and so that the lab can also track its performance.  SGS

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duplicate assays of pulps show good correlation as illustrated in Figure 14.1 and assays of standards as shown in Figure 14.2 are consistent and well within acceptable ranges of error.

Routine assay checks with standards (produced from mine rock taken from the mill circuit) are conducted at all times.  All standards are certified by SGS.  In 2004, Goldcorp started to make additional standards at Geoscience Laboratories, because of the recent purchase of Lakefield by SGS and to have access to another facility.  The mine’s normal practice is to insert two standards and three blanks per 100 samples collected.  There are four different types of blanks used and these are also created from mine material.  In general, standards in the economic grade range tend to assay slightly below the established grade, with occasional spikes above and below.  Four standards are now in use at the mine representing low, intermediate and High Grades and the labs are routinely tested for accuracy and precision.

14.2        Campbell Complex

Upon being drilled, core is transported to the core facility on the property.  After logging an sampling are completed, core samples are delivered to the on-site assay lab.  Chip samples are taken directly to the assay lab by the production geologist upon reaching surface.  Both whole and cut core samples are taken over a 6” – 3.0 ft (0.15- 1.0 m) interval depending upon the width of the structure.

14.2.1     Analysis (CC)

All assays are generated using our in-house assay laboratory. A constant monitoring of QAQC practices/procedures are conducted on a regular basis to ensure low assay contamination and consistent data accuracy.  All assays are verified by the chief assayer.

The geology and mine department submit shift samples for gold assay. The samples are received wet in plastic bags. These samples are placed into clean metal pans; the tag numbers scanned into the receiving database and the samples are placed into the dryer.

The dry samples are crushed to –1/4” by the jaw crushers and split through the riffle splitter into two parts (Rerun and Reject +100gm minimum). The sample is pulverised to –100 mesh and placed with the tag in a mixing can on the sample tray. Pulverisers are cleaned with air between samples to eliminate contamination between samples. Samples must be handled under the dust hoods to reduce exposure to quartz and silica dusts. Visible gold samples are crushed and the whole sample is pulverised in the TM ring pulveriser then submitted for fire assay.

The trays of 30 samples are mixed for 15 minutes prior to pulp weighing for atomic absorption or fire assay.

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The Sly dust collector and Torit are cleaned daily at the end of each shift.

14.3        QA/QC Procedures (CC)

Drilling, sampling, analysis, data stewardship, orebody modeling, and mine planning are carried out in accordance with NI43-101 industry standards. Regular internal auditing of the mineral reserve and mineral resource estimation processes and procedures are conducted.  The sampling and analytical methods are believed to be appropriate for the style and type of mineralisation. Mine geological staff has verified all databases used to generate the geological models and mineral resource estimates.

All samples have barcode tags and are scanned into the computer at sample receiving. Wet samples are then dried to remove any moisture.

Samples are crushed in a one stage or two stages, depending on the size of the coarse material. The crushed material is passed through the riffle with a minimum of 100 grams is kept for the rerun and reject portions. The sample tag is split and stored with each portion. The reject is stored in the sample prep area for 3 days and then discarded to the waste bin (paper tags removed to domestic garbage). The rerun portion is pulverised in the disc or vibratory ring pulveriser. These units must be cleaned after each sample. The sample is placed in a mixing can on a tray and mixed for 15 minutes. The sample numbers are scanned to generate the position on each tray.

The pulps are assayed by atomic absorption on a 5 gram sample. If the assay is > 0.05 oz/ton on a core sample, a fire assay is done on the original sample and the reject is also assayed.

Each tray has a quartz blank, a duplicate assay and two quality control standards. If either, the blank assays >0.01 oz/t, or the standards are outside +/- 10% the certified mean this will require that tray has to be re-assayed.

In addition, geologists are responsible for reviewing assay results. If required reruns can be requested.

All pulps other than core are dumped at the end of each day. The containers are cleaned with compressed air. The core samples are stored in paper bags and placed in cardboard boxes.

Assay results are entered into the database and approved daily.

Fire assays on process samples (mill) are reported in grams/tonne.  Fire assays on U/G samples are reported in oz/ton.  The Conversion Factor for ounces per short ton to grams per metric tonne is to multiply the value in ounces per ton by 34.286 to get the number of grams per metric tonne.

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To achieve the silver and gold Assays the following is done:

•The weight of the silver inquart is recorded.

•Weigh and record the gold/silver bead after cupellation.

•Subtract the weight of the silver inquart from the assay bead, this is the net weight of precious metal.

•The weight of the gold is subtracted from the precious metal weight, this is the silver content.

•Calculate and report in oz/ton or gm/tonne.

15.0        DATA VERIFICATION

15.1        Red Lake Complex

As part of the annual process control the Red Lake Complex undergoes an independent audit of its Mineral Resources.  The 2005 WGM audit of the High Grade Zone database and geological interpretations included:

•reviewing and spot validating the Gemcom® database supplied by the Goldcorp mine staff;

•checking zone interpretations, solid models and digitized boundaries on cross sections and level plans;

•reviewing statistical analyses of the main zones to corroborate cutting/capping parameters;

•spot checking of zone identification, composite grades and horizontal width calculations and transfer of this information into appropriate polygons for grade and tonnage estimation;

•checking zone calculations and totals completed outside of Gemcom (all final tabulations are done using MSExcel spreadsheets); and,

•verifying classification and reporting of Mineral Resources and Reserves.

All data used for WGM’s Mineral Resource audit was supplied by Goldcorp.  WGM did not verify information from drill logs or assay certificates, generate any new data or interpretations or perform an independent sampling program.  WGM reviewed the QA/QC program and the logging and sampling/assaying procedures and concluded that the database was in good order and that the

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procedures were to industry standards.  WGM assumed that the data supplied by Goldcorp was correct and accepted it for the purpose of its 2005 report.

15.2        Campbell Complex

Data from drill logs are checked internally during entry using QC routines in acQuire® to check for gaps, overlaps and duplicate entries.  The data then runs through a final check when the logging is done and the data is set for approval.

Assays are checked prior to import.  They are always first imported as a check only.  This will flag any errors, so the errors can be rectified before the actual import into the database.

Datamine is used as a final check to verify the location and accuracy of chip samples and drillholes.

After all drillhole data is complete with logging, assays and surveys a final approval is done and the holes are locked.

There are many internal checks done in the acQuire database to verify the data along the way.

16.0        ADJACENT PROPERTIES

With the combining of the Red Lake Complex with the Campbell Mine Complex there are no longer any adjacent properties that have any significant bearing on this report.  Although all the surrounding land in the area is staked by both Red Lake Gold Mines and junior companies, at present there is no property that would have any impact on the operating Red Lake Gold Mines.

17.0        MINERAL PROCESSING AND METALLURGICAL TESTING

17.1        Red Lake Complex

The original mill was built in 1948 and was dismantled in early 2000.  Construction of a new mill took place during 2000.  The new process facilities consist of three separate plants: the Crushing Plant; Processing Plant; and Paste Fill Plant.  Commissioning of the Crushing Plant began in February 2000, the Processing Plant’s commissioning phase commenced in early July 2000 with the first gold bar being poured on August 1, 2000 and commissioning of the Paste Fill Plant began in August 2000.  Commercial production began on January 1, 2001.

The Crushing Plant is a two-stage process which reduces underground ore from roughly 12 inches to 3/8 inches.  Underground ore from a coarse ore bin is fed to the Jaw Crusher and sizing screen.

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Screen oversize is crushed in the Cone Crusher and screen undersize is conveyed into the fine ore bin as feed material for the Processing Plant for gold extraction.

Unit operations in the Processing Plant include grinding, gravity concentrating, cyanidation, carbon-in-pulp, carbon elution and reactivation, electrowinning, bullion smelting/refining, cyanide destruction, flotation, and concentrate handling.  Three types of gold occur in the Red Lake Mine ore requiring these various unit operations.

Coarse gold is recovered from the ore via the gravity concentrating circuit.  A portion of the ground slurry from the single 1200 H.P. ball mill is fed to two Knelson Concentrators which produce a gravity concentrate that is upgraded on a Diester Table to a concentration of approximately 75% gold, and directly smelted into bullion.  The bullion is then shipped to the refinery for later sale into the spot market.  During 2005, the gravity circuit recovered 55% (345,585 ounces) of the gold from the processing plant feed.

Finer grain gold is dissolved in the cyanidation or leach circuit in which sodium cyanide is introduced to the process stream.  The leach circuit consists of four tanks each overflowing from one to the next.  In the leach tanks the gold is dissolved from a solid state into solution.  Gold is removed from solution and onto granular carbon particles in the C.I.P. (Carbon in Pulp) tanks.  Values from the carbon are removed in the Carbon Strip Plant, in which a high grade gold bearing solution (loaded eluate) is generated.  This loaded eluate, or pregnant solution, reports to two electrowinning cells where, under an applied voltage and current density, gold precipitates out of solution and back into its solid state as “cathode sludge”.  This sludge is also directly smelted into bullion for subsequent shipment to the refinery.  During 2005, 34% (218,778 ounces) of the gold contained in the Processing Plant feed was recovered in the Cyanidation Circuit.

The pulp discharging from the C.I.P. circuit is pumped to the Detox or INCO SO₂ circuit for cyanide destruction.  The circuit consists of two tanks with mechanical agitation where air, copper sulphate and sulphur dioxide are added to rapidly oxidize the cyanide and convert it to a non toxic cyanate that hydrolyses to ammonia.  The slurry feed to the Detox tanks contains approximately 200 ppm cyanide and discharges at less than 1 ppm cyanide.

The refractory component of the ore is gold that is extremely fine and locked in arsenopyrite and pyrite minerals (sulphides).  During 2005, 9% (55,969 ounces) of the gold in the Processing Plant feed was contained in the sulphide concentrate.  Conventional milling methods are not capable of recovering this type of gold.  The Red Lake Mine’s Processing Plant employs a typical sulphide flotation circuit generating a bulk sulphide concentrate.  This concentrate is subjected to further treatment for gold extraction on a custom processing basis.  It is shipped to either Barrick Gold’s Goldstrike Mine or the Campbell Mill autoclave.  From August 1, 2000 until December 31, 2005, 60,670 tons were shipped for a credit of 213,889 ounces of gold.  Concentrate that cannot be

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treated at the Campbell mill because there is not enough capacity is stockpiled for shipping to Goldstrike in the summer when above freezing temperatures allow reclaim of the thawed concentrate.

The process stream (tailings) reports to the Paste Fill Plant where most of the water is removed and the pulp is stored in a large stock tank.  This material is either discharged to the Tailings Management Area or sent underground for use as backfill.  The Paste Fill Plant is a semi-batch process, which implies that all aspects of the plant are continuous with the exception of the discharge of paste to the Underground Distribution System.  In the Paste Fill Plant a tailings filter cake is generated, cement and water is added and mixing occurs.  Once the proper consistency is achieved, the paste is discharged underground to flow by gravity to the mined out areas.  Figure 17.1 illustrates the Process circuit.

17.1.1     Red Lake Mill Performance and Limits

The Red Lake Complex mill treated more than 500,000 ounces on an annual basis from 2001 to 2005.  During 2001 and 2002 improvements were made in the flotation circuit to increase the plant recovery to the 97% range achieved in 2003 – 2005 which has also been maintained in 2006.  Figures 17.2 and 17.3 show the annual variation in feed ounces and recoveries.

Figure 17.2 Annual feed Ounces

Yearly Feed Ounces

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Figure 17.3 Annual Recovery Rates

Yearly Recovery

On a monthly basis the mill has seen feed ounces of over 70,000 ounces which equates to approximately 850,000 ounces on an annualized basis.  Normal recoveries of gold were achieved during the months in which the higher feed ounces were fed to the mill.  It is felt that the mill could handle feed ounces of up to 1,000,000 per year without a significant impact on recovery.  However these ounces would have to be delivered in the form of higher feed grades because the plant is limited on tonnage capacity.

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Monthly Feed Ounces

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Figure 17.4 Comparison of Recovery Rates against Feed Ounces

Monthly Recovery VS Feed Ounces Aug 2002 - Jan 2006

The mill was designed to process no more than 700 tonnes per day.  Levels of over 750 tonnes per day have been achieved with no noticeable affect on gold recovery (See Figure 17.5).  At levels above 800 tonnes per day negative impacts were observed on both coarser grind and a resultant lower gold recovery (Figure 17.5).

Figure 17.5 Comparison of Grinding Tonnage to Recovery Rates

Recovery VS Grinding Tonnage

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In 2005 an expansion project was started which is scheduled for completion the first quarter of 2007.  This project consists of a reclaim facility to be able to receive the ore from the No. 3 shaft, a new vertimill and upgrades to pumps and tanks within the mill.  With the addition of the vertimill and additional upgrades the new mill plant capacity will be greater than 1,200 tonnes per day.

17.2        Campbell Complex

The mill was designed to treat free milling and refractory gold ore at a rate of 360 tonnes per day in 1949.  The throughput has been gradually increased over the years to the current 1,850 tonnes per day.

Figure 17.6 shows the simplified mill flow sheet.  Conventional crushing and grinding is followed by gravity concentration to recover free milling gold.  Refractory gold, finely disseminated in the arsenopyrite and pyrite matrix, is recovered by flotation followed by pressure oxidation, neutralization and CIL. This steam joins the non-refractory flotation tails and is recovered by cyanidation/CIP processing.

17.2.1     Crushing

The ore is hoisted from the Reid Shaft to a 1,500 coarse ore bin. From there, it is transferred to a 250 tonne coarse ore bin located in the Campbell Mine head frame. The crushing plant consists of two Ross feeders, 457-mm x 915-mm (18-in x 36-in) jaw crusher,
1.3-m (4¼-ft) standard cone crusher, 1.3-m (4¼-ft) short head cone crusher, 2.1-m x 3.0-m (7-ft x 10-ft) Tyler double deck screen, variable speed short feeder belt, 6 conveyors.  A 19-mm product is produced in three stages of crushing at an average rate of 140 tonne per hour.  A jaw and standard cone crusher operates in open circuit and a short head cone crusher operates in closed circuit with an
18-mm vibrating screen.  The closed side setting of the standard and short head cone crusher is approximately 19-mm and 15-mm, respectively.

Fine ore is conveyed to the mill by inclined conveyor discharging, via a conveyor, to a 3,100-tonne fine ore bin.

17.2.2     Grinding and Gravity Circuit

Grinding is achieved in a two-stage rod/ball mill circuit.  The ore from the fine ore bin is fed to the rod mill via two slot feeders and a conveyor.  The grinding circuit consists of a 2.74-m x 3.8-m (9-ft x 12.5-ft) rod mill and 3.8-m x 4.7-m (12.5-ft x 15.5-ft) ball mill discharging, through trommel screens, into a common primary pump box.  The slurry is pumped to a cyclo-pac with 8 units of 250-mm CVX10 cyclone.  The cyclone overflow and underflow

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report to the flotation and ball mill, respectively.  Two cyclones, one feeding each Knelson concentrator, are mounted on independent underflow boxes away from the cyclo-pac.  These boxes are equipped with a concentrator feed inlet and an overflow return line to the primary pump box.  The concentrator cyclones are fed from the cyclo-pac distribution manifold and the overflow returning to the cyclo-pac overflow launder.

The grinding circuit produces flotation feed with an average p80 size of 73 microns (81% passing 200 mesh) and pulp density of 35 % solids by weight.  Mill throughput is dependent on underground ore supply so wide fluctuations in tonnage from day to day are possible, averaging 1,200 tonnes per day. The mill has a maximum milling capacity of 1,850 tonne per day.

The rod and ball mill is operated at 26.7 rpm, 67% of critical speed, and 22 rpm, 74% of critical, respectively. The rod mill is powered by 500 horsepower motor while the ball mill has a 1250 horsepower motor.

Shaking table concentration is carried out on the Knelson concentrate.  The final gravity concentrate assaying 72% gold by weight is refined into bullion.

17.2.3     Flotation Circuit

The cyclone overflow is gravitated to a 3-m² Delkor screen and the undersize to a pump box that is then pumped to a 10-minute conditioner tank.  The slurry reports to a 7-cell bank of Denver DR-500 rougher cells.  These cells are designed with double transverse froth overflow launders per cell and discharge into a single froth launder.  Concentrate reports to a 4-cell bank of Denver DR-100 cleaner cells.  Cleaner tails are recycled back to the DR-500 cells and the final concentrate assays approximately 15% sulphur and is pumped to a 9-m concentrate thickener.  The overflow from the concentrate thickener is recycled to the conditioner.  The flotation tailing is transferred to a 27-m diameter thickener with the underflow sent to the flotation tails leaching circuit and the overflow to the process water tank.  The reagents are stage added to the conditioner and junction box.  A Courier 30XP on-stream analyzer is used to monitor and control the flotation performance.

17.2.4     Pressure Oxidation Circuit

The pressure oxidation circuit, replaced the roaster circuit in July 1991, was designed to treat 71 tonnes per day of flotation concentrate or approximately 12.7 tonnes per day of sulphide sulphur. Carbonate destruction prior to pressure oxidation improves the oxygen utilization in the autoclave.  The thickened flotation concentrate (at 55% solids) is contacted with acidic solution (recycled 1^st CCD wash thickener overflow) in the pre-treatment circuit consisting of five 3.7-m x 7.3-m and one 3-m x 3.7-m pre-treatment tanks with a total retention time of 6 hours.  The

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recycled acid is generated by the oxidation of sulphides and reacts with the carbonates in the concentrate, evolving carbon dioxide. These pre-treatment tanks were equipped with 60 horsepower Lightnin A-310 agitators to enhance the reaction kinetic between acid and carbonates.  These agitators produce a pumping capacity of 150,000 gallons or eight tank turnovers per minute.  Overflow launders were installed between the six tanks with upcomers to minimize short-circuiting.  Fresh acid, 93% concentration by weight, can be added as required to maintain a discharge pH of not higher than 3.0.

The pre-treated slurry is transferred to an 11-m thickener with the overflow reporting to the waste treatment circuit.  The underflow is mixed with recycled 1^st wash thickener underflow before being pumped to the autoclave.  The recycling of solids provides a heat sink for the exothermic heat of oxidation to assist in temperature control and prevents the agglomeration of elemental sulphur.  Pressure oxidation is carried out in a five-compartment autoclave equipped with 50 horsepower Lightnin radial flow agitators, the first large compartment having two agitators.  The autoclave is lined with one layer of 6-mm lead and two courses (64-mm per course) acid resistant brick. The autoclave is 2.8-m in diameter with a tangent-to-tangent length of 15.2-m.  The autoclave operates at 190^oC and 2,100 kPa with a design slurry retention time of 2 hours.  VSA oxygen, at 93% in purity, is bought over the fence and injected beneath each agitator except in the last compartment (#5 agitator).  At a concentrate grade of 13% sulphur, the heat generated from the oxidation of sulphide minerals is more than sufficient to sustain the reaction and quench water is added to each compartment to provide further temperature control.  During the start up, a total of 4,000 kg per hour of steam is injected at 2,480 kPa into each compartment to bring the autoclave temperature to 190^oC (25^oC per hour).

The slurry within the autoclave cascades from compartment to compartment.  The level is controlled in the last compartment by regulating slurry discharge to an atmospheric pressure brick lined flash tower through a ceramic choke.  Flashing of steam reduces the slurry temperature to about 100^oC.  The slurry flows by gravity into a seal tank into which 2^nd wash thickener overflow is added to control the slurry temperature to 75^oC.  The slurry is then pumped to a two-stage counter-current decantation (CCD) wash circuit. The overflow from the 1^st wash thickener is recycled to the pre-treatment circuit previously described.  The underflow is split, with a portion being recycled to the autoclave feed and the reminder pumped to the 2^nd wash thickener for washing with fresh water.  The 2^nd wash underflow is neutralized with lime and transferred to an oxide carbon in leach circuit.

Cyanidation and carbon adsorption of the oxidised concentrate takes place in two 7.3-m x 7.3-m high tank with a retention time of 48 hours each.  The slurry is in contact with carbon at a concentration of 35 g/l. The leaching and carbon adsorption are not completed in this single stage circuit therefore the tails of the second CIL tank is combined with the flotation leach and CIP circuit.

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17.2.5     Flotation Tails Leaching and Carbon-In-Pulp Circuit

Thickened flotation tailing (60% solids) is leached for 20 to 28 hours.  The initial pH and cyanide concentration is maintained at 11.5 and 500 ppm, respectively.  The total lime and cyanide consumption on both oxide and flotation tailing circuit is 2.1 and 0.75 kg/tonne.  The leached slurry, a combination of oxide and flotation tails, is pumped into a train of six CIP tanks; each has a slurry retention time of 50 minutes.  Carbon retention is accomplished by the use of 2.25 m² KEMIX wedge screens with a carbon retention time of approximately 10 days (carbon concentration 35 g/l).  The 6 x 16 carbon advances 8 times over a 24-hour period counter-current to the pulp flow by a SALA vertical recessed impeller pump suspended in the pulp.

Carbon is transferred from the CIP #1 tank to the CIL tank to increase the carbon grade to approximately 9,000 g/t. Approximately 1.3 tonne per day of loaded carbon is removed every day in a single operation with a SALA vertical recessed impeller pump that discharges over a 20-mesh screen.  The carbon flows by gravity into a loaded carbon bin and undersize is returned to the CIL tank.  Acid washing is performed using 5% by weight hydrochloric acid on every second batch of loaded carbon.  The loaded carbon is stripped using 40 bed volumes of 1% caustic solution and 0.1 % cyanide at 140^oC and 480 kPa.  Stripped carbon is educted from the strip vessel over a 16 mesh de-watering screen.  The screen oversize (stripped carbon) flows into the kiln feed hopper for de-watering and undersize (fine carbon) into the eductor water tank.  The de-watered stripped carbon is then fed into a carbon regeneration kiln with zone 1 and 2 set at 450 and 650^oC, respectively.  The stripped solution is pumped to the electrowinning cell for gold plating. The barren solution is recycled to the strip solution tank.

17.2.6     Refinery

The gravity concentrate and electrowinning sludge are melted simultaneously. The mixture of gravity concentrate and EW sludge is fluxed with nitrate, borax, silica, fluorspar and manganese dioxide.  The charge is melted in a 4.5 cu ft Bradley induction furnace.  The reagents combine with impurities to form a slag that is recycled.  The gold is molten, purified and poured into 1,000 troy ounce bullion bars.  The bullion averages 92% gold and 6% silver.

17.2.7     Paste fill and Waste Treatment Circuit

The CIP tails are partially deslimed through a set of cyclones with a mass flow split of 90-10%. The underflow is sent directly to the paste thickener.  From there, it is pumped to two disc filters and mixed with cement and flyash to form a paste. The paste is pumped underground via a high-pressure piston pump. The cyclone overflow combines with acidic overflow from the pretreatment thickener in the waste treatment circuit consisting a series of four agitated tanks.  Lime is added to tank #3 to control the final discharge pH that is set at 8.5 to 9.0.  At this pH, the

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formerly complexed metals precipitate out along with the other dissolved metals as hydroxides.  The final tailings discharge to the main tailings pond.

17.2.8     Effluent Treatment Circuit

The effluent treatment circuit consists of two reaction tanks, a clarifier feed tank and a 9.5-m diameter hopper (double-vee) clarifier.  Between May and December each year, decant from the main tailings pond is pumped back to the mill for cyanide destruction and metals precipitation. Cyanide destruction is accomplished using the Inco SO2/Air process in agitated and aerated tanks. Cyanide is destroyed in tank #1 at a pH of 9.0 and metals, primarily copper and nickel, are precipitated in tank #2 at a pH that varies from 9.5 to 11.0.  The treated solution is then transferred to a hopper clarifier with the overflow reporting to the settling pond. The sludge recovered from the bottom of the cone is partly recycled to the clarifier feed tank and the remaining to the waste treatment circuit.

17.2.9     Polishing Pond and wetland

The polishing pond, commissioned in November 1995, consists of a 400,000-m³ settling pond and 730,000-m³ holding pond.  A centre dyke separates these ponds.  At a feed rate of 15,000-m³ per day, the ponds have a retention time of 75 days.  The hopper clarifier overflow discharges into the settling pond where the residual ultra-fine precipitate (complexed hydroxides) are settled before progressing to the holding pond.  The pond is operated on a seasonal basis.  During the warmer months, the water level is allowed to rise to the operating level at which time the discharge rate from the polishing pond is matched to the inflows to the pond.  Natural degradation in the holding pond improves the quality of the water.

Since 2001, the discharge of the polishing pond is directed to a series of cells heavily vegetated with cattails. The water residence time is approximately 2 to 3 days in the wetland.  The effluent quality is further improved with an 85% reduction in ammonia and 50% reduction in copper concentrations. The effluent meets all acute toxicity tests for rainbow trout and daphnia magna.

17.2.10  Planned Flowsheet Developments

Capital projects completed in 2005 were directed at increasing gold recovery through modifications to the autoclave weirs and agitators. A larger electrowinning cell was also installed.

Capital projects completed in 2006 were directed at increasing tonnage through the autoclave. Several iterations were investigated to supply the plant with oxygen and upgrades are to be completed by Q-1 2007.

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17.2.11  Production Performance

Recent gold production has averaged 200,000 ounces per year. Gold recovery is strongly affected by head grade and has remained above 95% the past 5 years. Mill utilization has been poor due to variable mine supply and the need to produce paste fill to maintain mining cycles. Daily throughput shows the mill can easily perform at 1,500 tpd but has recently performed well at above 1,850 tpd for sustained periods.

Figure 17.7 Monthly Gold Production for the Campbell Complex

Figure 17.8 Monthly Mill Recovery for the Campbell Complex

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Figure 17.9 Monthly Mill Utility for the Campbell Complex

Figure 17.10 Average Daily Mill tonnes by Month for the Campbell Complex

18.0        MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

The mineral estimate was based on two methods of calculation.  At the Red Lake Complex mineral estimate was completed using longitudinal polygonal techniques using Gemcom® geological software system.  For the purposes of this report, the estimate has not been updated since the end of 2005.

At the Campbell Complex 3D geological and block models were constructed with Datamine® mine planning software based on the September 2006 database.

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The reserves are based on a US$450 /oz gold price and exchange rate of 1.15.

18.1        Red Lake Complex

The 2005 Mineral Reserve estimates were completed by Red Lake Mine and audited by Watts, Griffis and McOuat Limited (WGM).  The following section 18.1 has been taken from that report.

Goldcorp is listed on both the Toronto and New York Stock Exchanges and must satisfy the requirements of both jurisdictions.  For the purpose of the WGM audit and report the definitions for Resource and Reserve have been guided by the SEC definitions for Proven and Probable Reserves.

For the purpose of current disclosure rules in Canada following the guidelines of National Instrument 43-101 (“NI 43-101”), WGM confirm that the Proven and Probable definitions used by the SEC conform with USGS Circular 831, which is acceptable under NI 43-101.  WGM further confirm that these definitions and the Reserve numbers can be reconciled directly with the CIM Standards, as reported in NI 43-101, without further adjustment.

The Red Lake Mine contains three types of mineralisation; vein type, disseminated sulphide and replacement ore.  Mineral Reserve estimates were completed at a US$400 per ounce gold price for both the High Grade and Sulphide Zones for 2005.  The High Grade Zone mineralisation is defined at a cutoff grade of 0.40 opt and the Sulphide Zones are defined at an approximate cutoff of 0.23 opt.

There are 18 sub-zones with Proven Mineral Reserves presently defined within the High Grade Zone, and six of these zones (EW, FWA, HW, HW5, HWA_2 and MAIN_1L) account for about 73% of these tons and 76% of the contained ounces (Table 18.1).

Please note that mineral resources that are not mineral reserves do not have demonstrated economic viability.

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Table 18.1     Red Lake Complex Reserves (as of September 30, 2006)

Area/Category	Tons	Grade (oz Au/ton)	Contained ounces of Gold
High Grade Zone
Proven (incl. Broken ore)	677,000	2.27	1,538,000
Probable	1,119,000	2.53	2,831,000
Subtotal	1,796,000	2.43	4,369,000
Sulphide Zone
Proven	204,000	0.40	82,000
Probable	1,122,000	0.41	465,000
Subtotal	1,326,000	0.41	547,000
Mine Wide, All Zones	881,000	1.84	1,620,000
Proven (incl. Broken ore)	2,241,000	1.47	3,296,000
Probable	3,122,000	1.57	4,916,000
January 1-September 30, 2006
Mined from Reserve	183,449	2.09	383,921
Total Reserves	2,938,551	1.54	4,532,079

Note: Based on a gold price of US$400/oz and an exchange rate of US$1.00 = C$1.20.

In addition to the Proven and Probable Mineral Reserves, there are Remaining Mineral Resources in the Measured and Indicated category, which have not as yet been included in the mining plan for either the High Grade or Sulphide Zones, plus Inferred Mineral Resources based on more widely spaced drilling.  These Mineral Resources are as shown in Table 18.2.

Table 18.2     Red Lake Mine remaining Mineral Resources (as of December 31, 2005)

Area/Category	Tons	Grade (oz Au/ton)	Contained ounces of Gold
High Grade Zone
Measured	2,000	3.50	7,000
Indicated	306,000	0.87	265,000
Subtotal	308,000	0.88	272,000
Sulphide Zone
Measured	333,000	0.32	108,000
Indicated	1,601,000	0.34	547,000
Subtotal	1,934,000	0.34	655,000
Total Measured and Indicated	2,242,000	0.41	927,000
High Grade Zone Inferred	742,000	1.37	1,017,000
Sulphide Zone Inferred	1,015,000	0.54	549,000
Total Inferred Mineral Resources	1,757,000	0.89	1,566,000

Note: Based on a gold price of US$450/oz and an exchange rate of US$1.00 = C$1.20.

The Mineral Resources and Reserves of the High Grade Zone were estimated for the original Feasibility Study in September 1998, and subsequent year-end reports by Goldcorp to the present.

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In 2000, the ore zone configurations were modified as a result of structural studies and exposure by mining.  The cutting factor was revised upward to Mean plus 1 Standard Deviation (“SD”), or 15 g Au/t, which increased the grade from 1.38 oz/ton to 1.68 oz/ton.  Mining during 2001 produced grades exceeding the 2000 Mineral Reserve estimate and in August the cutting factor was increased to Mean plus 3 SDs on a per zone basis.

The 2002 mining results indicate that a slightly higher cutting factor should be used and Mean plus 4 SDs (per zone) produces the best comparison between production and estimated Mineral Reserves.  The cutting remained at this level through 2004, but was again revised for 2005.  Additional in-fill results have been incorporated in the 2005 Mineral Reserves and additional Mineral Resources and Reserves have been added to the down-plunge estimation of the HW and FW zones.

The modelling of the High Grade Zone and estimation of the Mineral Resources and Reserves were done by Goldcorp mine staff using longitudinal polygonal techniques using Gemcom® geological software system.

Traditionally, the Mineral Resources and Reserves of the Sulphide Zones above the 30 Level have been reported as they were estimated as of the end of 1997 by mine staff using manual methods.  Recently, Goldcorp has been concentrating on updating these Sulphide Zones in preparation for future production.  The E Zone between the 22 and 26 Levels, and parts of F and ESC Zones, and the Sulphide Zones below the 30 Level, were estimated at the end of 1998 by computerized polygonal methods as employed for the High Grade Zone.  During 2003, a major revision and re-interpretation of the Sulphide Zones incorporating new drilling information was being carried out.  This was continued into 2004 and is still in progress.  The Sulphide Zones that received the majority of the drilling and new modelling for 2005 were the Deep Sulphide, ESC3, Lower E, F Main and Far East zones.

18.2        Method of Resource evaluation (RLC)

18.2.1     Database and Validation

All digital data for WGM’s audit was supplied in Gemcom® format in one project database from the Red Lake Mine staff.  The project database contains four “Gemcom Workspaces” for drillhole, chip sample, polygon and proposed drillhole information.  The drillhole workspace contains over 21,000 collar location records with related tables for downhole surveys, lithology, mineralogy, alteration, structure, raw gold assays and various composite gold assay tables.

The polygon workspace contains records for grade and resource polygons and stope outlines that represent approximately 10,000 drillhole sampled or mined out areas.  The polygons in the

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database have a 3-D location and contain information such as zone and drillhole identifiers, cut and uncut composite zone gold grades, horizontal widths and areas.  The chip sample workspace contains over 22,000 records and has related location, geological, assay and other pertinent information.

WGM reviewed and validated selected parts of the drillhole database in order to identify such errors as overlapping or out of sequence intervals, missing assays, duplicate samples, etc.  Minor errors were found in the assay and composite tables, however, as they are not located in the defined ore zones, the errors do not effect the Mineral Resource/Reserve estimate.  WGM spot checked the drillhole database intervals and compared them to the polygon intervals to ensure that the relevant data for the zone was transferred properly from one workspace to the other.  The conversion of core length to horizontal width was also spot checked and viewed on cross sections.  We also compared the polygonal information contained in the database against the longitudinal plots used for the Mineral Reserve estimate.  WGM found no errors.  In general, WGM found the database to be accurate and in good order.

18.2.2     Geological Controls

A geological study of the Red Lake Mine was done in 1999 and a structural model was developed to explain the distribution of mineralisation in the High Grade Zone.  The basic concept is that of northwest shearing and northerly-trending dilatant fractures.  Underground development has shown that the mineralisation is somewhat more irregular than expected with a variety of vein attitudes on a small scale.

The HW Zone is the most conformable and considered to be the principal shear structure.  The zone has the highest ratio of drillhole intervals reported as carbonate veins rather than quartz vein replacement type, and has the largest vertical extent.  Of the five largest veins, the HW Zone is the lowest grade, which is attributed to its tight shear character.

The MAIN Zone is interpreted as a hybrid consisting of an off-set shear with a dilatant connection in the centre.  The uncut Mineral Resource grades increase from the shear type HW Zone, presently at about 1.5 oz Au/ton, through the MAIN Zone at about 2.5 oz Au/ton, to the most open vein type structure of the HW5 Zone, currently grading over 3 oz Au/ton.  Thickness trends in the main zones plunge most commonly at 45% to the northwest, but some of the small zones cannot be positively correlated until they are mined.  Detailed drilling and underground development has resulted in some of the zones being correlated differently and being revised from time to time.  Some revisions have been made to the Mineral Resources in the various zones, but the changes to the estimates over time have been relatively minor, except for expansion at depth of some of the zones.

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18.2.3     Zone Interpretation and Modelling

To prepare the Mineral Resource estimate, mineralized intercepts were selected based on geological features as well as consistent geometry and continuity of grade.  The method employed to define resources is similar for both the High Grade and Sulphide Zones.  As mining and underground mapping progresses, Goldcorp is gaining more experience with the mineralisation and geometry of the High Grade Zone and is continually refining its interpretations of the specific ore shoots.  Goldcorp is now more confident to include some of the lower grade fringe material that appears to be part of the structure, as this material is often turning out to be ore.  Recently, the MAIN and HWA zones have been broken up into high angle and low angle portions.  The experience with mining the High Grade Zone has also helped Goldcorp understand the geology of the Sulphide Zones better and is aiding in the reinterpretation of these zones.

The mineralized zones are first interpreted on cross sections at 25 (8m) and/or 50 ft (16m) intervals (Figure 18.1).  The Mineral Resource outlines are then projected on level plans to ensure the horizontal continuity of each mineralized shoot.  Figure 18.2 is a plan map showing the High Grade Zone mineralized shoots in red and the Sulphide Zones in orange.  The mineralized drillhole intersections are grouped by zones and projected on a series of vertical longitudinal projections sub-parallel to the general trend (Figures 18.3 and 18.4).  When no mineralisation was identified in a drillhole, a minimum interval was selected and plotted on the longitudinal projections.  Zone outlines were interpreted on each longitudinal projection using a general cutoff of 0.40 oz Au/ton (14g/tonne).  Locally, lower grade intercepts have been included in the zone outlines to form reasonable mining units.  This methodology takes into account the irregular nature of the mineralisation and produces a workable model for mine planning. The major lenses show good continuity along strike and dip.  The strike extension is typically about 200 ft (60m) and locally up to 400 ft (120m).  The down-dip extent of several individual zones has been traced for over 1,200 ft (365m), with the HW5 Zone now exceeding 3,000 ft (915m) vertically.

The average thickness of all of the sub-zones in the High Grade Zone is approximately 11 ft (3m) and maximum width is about 20 ft (6m).  For the Sulphide Zones, the average width is about 12 ft (3.5m) with a max. 18 ft (5m).  The average width is slightly larger below 30 Level than above, however, presently there are far fewer Sulphide sub-zones defined at depth.

18.2.4     Polygonal Mineral Resource Modelling

Based on the selected intercepts and on the Mineral Resource outline described above, a polygonal model on vertical longitudinal projections of each zone is created to estimate the tons and grade (see Figures 18.3 and 18.4).

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The maximum radius used for a polygon is normally 50 ft (15m).  Where the distance between two drillholes is less than 100 ft (30m), the limits of polygons are drawn by tracing a line at half of the distance between each pair of intercepts.  The shape of the polygons is further controlled by the geological outline interpreted for each zone.  These outlines are constantly reviewed and changed as required, as additional drilling information becomes available or if a drillhole intersection changes from one zone to another due to a revised geological interpretation.

The estimating parameters for 2005 year-end Mineral Resources and Reserves have remained essentially the same as last year, i.e., the continuing use of more “bounded cases” for many of the High Grade Zone sub-zones where it is deemed more practical and the deletion of “In-fill Inferred” polygons.  Over the last two years, most of the important zones have been reworked and reinterpreted in longitudinal view and the bounding case allows for a stronger geological influence to be applied on the shape of the polygons in longitudinal view, rather than just accepting the shape of the polygons without bounding, as was done previously.  Bounding a polygon model restricts the influence of the polygons to within the bounding outline and removes the influence of waste intercepts which fall outside of the outline.  The bounding outline was shaped and finalized based on compiled geological data, both drilling and production sampling.

The reinterpretation of these zones gave Goldcorp the opportunity to re-assess the Mineral Resources that were previously classified as In-fill Inferred.  Many of these “polygons” occurred within zones with very strong confidence of continuity (HW5 and HWA primarily) and therefore could justifiably be re-categorized as Measured or Indicated, even though the radius may have exceeded the standard 50 ft (15m).  The maximum radius used for a polygon was 50 ft (15m) in unbound sub-zones, however, in the newer bound cases, the radius is constrained by the geological interpretation in longitudinal view, but could potentially extend up to 150 ft (45m).  This means that the “In-fill” Inferred Mineral Resource polygons can now be substituted with larger continuous polygons.  This was done on a zone by zone basis and was only applied to polygons contained within zones with demonstrated strong continuity.  This more “hands-on” approach allowed Goldcorp to still categorize some larger polygons as Inferred Resources in selected areas where information and data density was not as high.

The bounded case has the effect of “adding” tons and contained ounces in the Measured and Indicated “Remaining Mineral Resources” classification, however, it decreases the total amount of Inferred Mineral Resources (so in effect, it is primarily a re-categorization of these Resources).  WGM is pleased with this development, as we have recommended moving to this method over the last couple of years.  This should improve the interpretation of the zones, thereby increasing the confidence of the estimate.  This new method has helped close the gap between the predicted Mineral Reserve and the actual production of ounces in areas already mined.

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The Mineral Resource tonnage associated with each polygon is calculated from the area of the polygon, the horizontal thickness of the intercept and the tonnage factor.  Narrow zones are expanded to a minimum width of four feet using wallrock at zero grade for the High Grade Zone and six feet for the Sulphide Zones.  The larger minimum width used for the Sulphide Zones reflects the historical use of higher internal dilution and Goldcorp has elected to retain this width for the Mineral Resource estimates.

18.2.5     Statistical Analyses

Ore Zone Assays

As in previous years, Mineral Resource and Reserve estimates for the High Grade Zone are based only on drillhole results and no allowance has been made for samples taken underground which have limited influence and are of a different character.

Thirty-two sub-zones have been included in the Mineral Resource estimate for 2005.  This total changes from year to year as new zones are discovered or as larger zones are split into smaller sub-zones as Goldcorp gets a better understanding of the geometry of the zones from mining and/or additional drilling.  The data within the zones is extracted from the raw assay tables in the Gemcom® database.  The basic statistics on the assay intervals (Table 18.3) indicate the unique nature of the mineralisation, with approximately 38% of the assays above 1 oz Au/ton.  The population has a log normal distribution.

Table 18.3     Basic Statistics on High grade Zone Raw Assays

Uncut Assays	All Data	oz Au/ton
		<0.20	0.20<1.00	>1.00
Mean	4.50	0.06	0.50	11.29
Variance	572.66	0.00	0.05	1,410.85
Median	0.49	0.05	0.46	3.45
Minimum	0.001	0.001	0.200	1.000
Maximum	1,172.400	0.199	0.999	1,172.400
No. of Samples	12,790	4,663	3,198	4929
% of Samples	100.0	36.5	25.0	38.5

Normalized Samples

There are 12,790 raw assays inside the Mineral Resource outlines.  Before cutting the high values, the data is normalized to equal lengths of two feet, based on the approximate average sample length within the zones.  This brings the number of samples to 13,353, as sample remnants (i.e., less than two feet) are also included for grade estimation purposes.  These normalized samples are then used to calculate the composite grade of the entire zone.

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Basic statistics on the two-foot normalized samples are presented in Table 18.4.  The proportion of samples above 1 oz Au/ton (34g/tonne) still remains high at almost 40% of the total population.

Table 18.4     Statistics on High Grade Zone two-Foot (60cms) normalized assays

UNCUT ASSAYS

Normalized Uncut Assays	All Data	oz Au/ton
		<0.20	>0.20<1.00	>1.00
Mean	3.63	0.07	0.51	8.74
Variance	211.13	0.00	0.05	489.12
Median	0.58	0.05	0.47	3.14
Minimum	0.000	0.000	0.200	1.000
Maximum	588.562	0.199	0.9995	588.562
No. of Samples	13,353	4,340	3,718	5,295
% of Samples	100.0	32.5	27.8	39.7

Note: “Remnant” samples (i.e., mineralisation less than 2 ft) are included in the statistical analyses and cutting procedures.

18.2.6     Cutting and Capping Procedures

In the pre-production Mineral Resource estimates, the historical 2-5-10 oz Au/ton (69-171-343 g/tonne) cutting procedure was used for both the low grade Sulphide Zone and the High Grade Zone.  The 2-5-10 rule which involved reducing very high assays to 10 oz Au/ton, 5 to 10 oz Au/ton to 5 oz Au/ton, and 2 to 5 oz Au/ton, to 2 oz Au/ton produced a Mineral Reserve grade significantly lower than the average mill head for the first 92,800 tons mined in 2000.  This was an expected result, since the historic cutting procedure was based on the much lower grade sulphide ore.

For the 2000 year-end Mineral Reserves, a top cut of Mean plus 1 SD (15 oz Au/ton or 514g/tonne) was adopted in order to estimate a reserve comparable to the production results that were being obtained at that time.  Due to complications caused by excessive dilution during development and the small tonnage involved, the cutting factor could not be determined with proper accuracy and subsequent production results continued to exceed predictions.

Results to August 2001 indicated that a further revision to the cutting factor was required to reconcile the Mineral Reserve estimate more closely with the mill head grade.  The cutting factor was again adjusted upward, this time to Mean plus 3 SDs on a per sub-zone basis.  At the end of 2002, mill head grade continued to exceed Mineral Reserve estimates, and the cutting factor was revised to Mean plus 4 SDs.  This capping level was retained until this year, as an error was identified in the method that Goldcorp used to calculate horizontal widths for certain sub-zones.  This error was significant to only a small group of mostly east-west oriented drillholes and was corrected in 2005.  The error resulted in narrower horizontal widths being reported and used for

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the Mineral Resource estimate than should have been, which ultimately meant that fewer ounces were being reported.

In order to reconcile the estimates with production for these affected sub-zones, Goldcorp could not use the previous Mean plus 4SD cut, as it would produce more ounces in the total Mineral Resources, and too many ounces in the reconciliation calculation.  To better reconcile with production, Goldcorp reviewed the last four years of production on a per sub-zone basis and decided to go back to the Mean plus 3SD for most of the sub-zones in the High Grade Zone, except for HW5, HW7 and HWA.  The statistics were recalculated with this revision and Table 18.5 shows the old (2004) and new capping levels for each sub-zone.

This correction increased the horizontal width in most of the holes affected and therefore resulted in an addition to the Mineral Resource estimate of about 70,000 contained ounces, or approximately 1.3% of the total High Grade Zone Mineral Resources.

Using individualised capping levels for the various sub-zones ensures greater accuracy because of the considerable difference in grade and characteristics of each population.  Some individual Sulphide Zone Mineral Resources and Reserves have been cut more aggressively than the traditional 2-5-10 cutting factor where data and experience warrant being more conservative.

Table 18.5     Per Zone Top Cuts on High Grade Zone Two-Foot (60cms) normalized samples (December, 2005)

	No. of		Standard	2005	2005	2004
Zone	Data	Mean	Deviation	SD	December	December
	Points		(SD)	Multiplier	Top Cuts	Top Cuts
EW	657	7.40	29.04	3	94.51	104.79
EW1	75	2.23	5.77	3	19.56	104.79
FW	305	6.63	19.46	3	65.02	58.69
FW2	166	1.51	4.61	3	15.32	14.41
FW3	188	0.74	1.38	3	4.88	5.49
FW3A	27	0.33	0.37	3	1.42	1.42
FW3B	153	1.81	5.16	3	17.29	26.27
FW4	781	0.85	2.17	3	7.34	5.90
FW4A	219	0.96	2.31	3	7.90	5.50
FW4B	143	1.74	4.10	3	14.04	9.24
FW4C	268	1.28	5.57	3	17.99	6.28
FWA	364	5.61	16.58	3	55.36	61.75
FWB	143	1.74	4.10	3	14.04	11.33
FWC	37	2.58	3.96	3	14.44	13.59
FWD	100	2.28	4.74	3	16.49	21.26
FWX	66	2.90	6.11	3	21.22	26.06
HW	1443	2.80	11.05	3	35.94	34.15
HW5	1802	6.07	22.77	4	97.16	76.43
HW5A	28	0.84	0.97	3	3.75	4.06

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	No. of		Standard	2005	2005	2004
Zone	Data	Mean	Deviation	SD	December	December
	Points		(SD)	Multiplier	Top Cuts	Top Cuts
HW5B	3	15.73	15.28	3	61.57	61.57
HW5C	87	4.15	12.93	3	42.95	4.12
HW6	44	1.01	2.02	3	7.06	34.15
HW7	63	1.18	1.53	4	7.28	44.70
HW7A	67	1.49	6.57	4	27.76	44.70
HWA_1L	288	1.96	9.99	4	41.91	11.72
HWA_2	487	6.42	19.02	4	82.48	97.14
HWB	97	1.75	4.69	3	15.82	10.22
HWD	731	1.71	4.11	3	14.06	15.77
HWX	121	3.83	10.93	3	36.62	32.82
HWX1	56	4.04	10.87	3	36.64	32.82
HWY	11	1.69	2.18	3	8.24	62.19
MAIN_1L	2186	2.30	7.27	3	24.12	25.92
MAIN_2	738	8.08	18.54	3	63.70	73.56
MAINA	962	2.22	5.82	3	19.68	21.50
MAINB	419	4.07	19.44	3	62.38	16.05
MAINX	90	3.99	15.68	3	51.01	50.20

18.2.7     Mineralized Intercepts

A total of 2,563 mineralized intercepts (i.e., normalized zone composites calculated over a horizontal width and imported back into Gemcom®) have been retained for the Mineral Resource estimate of the High Grade Zone.  The mean of the intervals is 3.92 oz Au/ton (134g/tonne).  Interestingly, the average projected horizontal width of those intercepts (number of two foot intervals divided by the number of zone intercepts x 2) is 10.4 ft (3.2m), which matches well with Goldcorp’s average mining width, as most stopes are 8-10 ft (2.5 – 3m) wide.  Table 18.6 shows the basic statistics of the mineralized intercepts using the uncut grade after calculating the grades using the two-foot samples.

Table 18.6     Basic Statistics on High Grade Zone uncut mineralized intercepts

	All Data	<0.20	>0.20<1.00	>1.00
		oz Au/ton	oz Au/ton	oz Au/ton
Mean	3.92	0.07	0.56	6.34
Variance	81.13	0.00	0.05	123.51
Median	1.31	0.04	0.54	2.79
Minimum	0.00	0.00	0.20	1.00
Maximum	112.32	0.198	0.998	112.32
No. of Samples	2,563	221	833	1,509
% of Samples	100.0	8.6	32.5	58.9

It is significant to note that nearly 60% of the intercept population exceeds 1.00 oz Au/ton (34 g/tonne).

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18.2.8     Grade Estimation, Cutoff and Mine Operating Costs

The grade of each polygon is the weighted average of the capped grade of two-foot normalized samples (60cms)  included in the intercept.  The dilution included to extend all intercepts to a minimum horizontal width of 4 ft (1.2m) (6 ft (1.8m) for the Sulphide Zones) is recorded as internal dilution and is a part of the Mineral Resource tonnage.  The dilution for both the Mineral Resources and Reserves (to a minimum of 7 ft (2.1m) for HGZ Reserves and generally a minimum of 9ft (2.7m) for the sulphides- see below) was calculated. Total grade of the zones was calculated by weighting the intersection grade by the tons represented by each polygon.

The geometry of the ore bodies and the capacity of the operation both contribute to a high cost structure for the Red Lake Complex and a relatively high cutoff grade.  However, the grade of the deposit allows Goldcorp to deliver a high mill head grade and to be a very low cost gold producer on a per ounce basis.  The economic cutoff for mining purposes is approximately 0.40 oz Au/ton (14g/tonne), including 3 ft (1m) of external dilution from the walls.

As outlined in “The Red Lake Mine Expansion Study”, prepared by Goldcorp in December 2002 (“the Expansion Study”), the planned expansion of the mill operation in conjunction with the development of the No. 3 Shaft will reduce the unit cost of operations and the cutoff grade.  The unit operating costs, including off-site processing, for 2005 are shown in Table 18.7.

Table 18.7     Goldcorp Unit Operating Costs - 2005

Operations Area	Cost per Tonne Milled (C$)
Definition	14.69
Development	6.86
Mining	138.99
Milling	30.07
Maintenance	36.90
General and Administrative	33.60
Stockpile Adjustment	-2.38
Off-site Concentrate Processing	16.72
Total Cost	275.45

Operating costs in 2005 totalled approximately C$275 per tonne milled.  At gold prices of about US$500 per ounce, or approximately C$600/oz per ounce, the break-even grade is approximately 0.47 oz Au/ton (16g/tonne).  The Expansion Study projects a unit cost of C$197 per ton for a break-even grade of 0.34 oz Au/ton (12g/tonne) at this gold price and with 97% recovery.

The present average grade of the Sulphide Zones Mineral Resource is identical to the cutoff grade of the High Grade Zone (i.e., 0.40 oz Au/t or 14g/tonne).  However, some material from the Sulphide Zones only qualifies as a Mineral Resource if it is evaluated from the standpoint of

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marginal additional revenue and marginal additional cost.  In its Expansion Study, Goldcorp prepared cash flow analyses based on two scenarios.  The first scenario used 800 tons per day from the High Grade Zone only, and the second scenario used 800 tons/day from the High Grade Zone plus 200 tons/day from the Sulphide Zones.  The latter scenario resulted in a higher rate of return.

In the opinion of WGM, this study supports classifying the Sulphide Zones material as Mineral Resources and Reserves.  This condition exists as long as the hoisting and processing capacity exceeds the production rate achievable from the High Grade Zone only.  If future exploration were to result in large increases in lower grade tonnage without commensurate increases in High Grade tonnage, further analysis would be required to define the economics of mining the Sulphide material.  Excess low-grade material would not be economic and could not be included in Mineral Reserves.  In 2005, the study was updated with a planned production rate of 1,250 tons per day with High Grade Zone and Sulphide Zone ore to be mined in the same 80:20 ratio.

18.3        Application of Mineral Resource Definitions (RLC)

Measured Mineral Resources are designated to specific “ore shoots” that have at least one mining excavation to confirm lateral continuity of the structure.  In addition, there must be sufficient drilling density on this structure above and below the excavation to demonstrate continuity on longitudinal sections.

Indicated Mineral Resources consist of two or more adjacent or almost adjacent polygons included inside a zone outline which are considered to be economically mineable.

Inferred Mineral Resources consist of isolated polygons or blocks that are not currently economic.  Previously, there was also an Inferred classification for the areas inside some of the High Grade Zone outlines where the polygons did not cover the interpreted zone completely.  This Inferred category was carried separately as “Inferred In-fill” in some zones, but is now replaced by using the bounded case and larger internal polygons where permitted.

18.4        Mineral Resource and Reserve Statement for High Grade Zone (RLC)

Whereas in 1999 Mineral Resources from nine sub-zones were classified in the Measured and Indicated category, this year the Mineral Resources in parts of 32 sub-zones have been classified in the Measured and Indicated category as a result of underground development and extensive in-fill drilling.  WGM believes that the upgraded classification is valid based on quality and abundance of data.

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18.4.1     Cut Grade Mineral Resource Estimate

The Measured, Indicated and Inferred Mineral Resources of the High Grade Zone using the cut grade (as per Table 18.5) are shown in Table 18.8.

Table 18.8     High Grade Zone Mineral Resources (Cut Grade) 2005

Category	Tons	Grade (oz Au/ton)	Contained ounces of Gold
Measured	469,000	3.25	1,525,000
Indicated	1,130,000	2.74	3,096,000
Total	1,599,000	2.89	4,621,000
Inferred	742,000	1.37	1,017,000

18.4.2     Uncut Grade of the Mineral Resource Estimate

In order to illustrate the level of the cutting procedure, the uncut grade of the Mineral Resource estimate is shown in Table 18.9.  For the Measured and Indicated Resources, the cutting procedure has reduced the uncut grade by about 20% and has reduced the contained ounces by a similar amount.

Table 18.9     High Grade Zone Mineral Resources (Uncut Grade) 2005

Category	Tons	Grade (oz Au/ton)	Contained Ounces of Gold
Measured	475,000	3.77	1,790,000
Indicated	1,130,000	3.57	4,033,000
Total	1,605,000	3.63	5,823,000
Inferred	742,000	1.41	1,046,000

The uncut Measured and Indicated Resources grade in 2005 at 3.63 oz Au/ton is somewhat higher than the 2004 grade of 3.23 oz Au/ton, due primarily to slightly higher grade results from in-fill drilling and the addition of higher grade Mineral Resources at depth in the HW5 Zone.

18.4.3     Mineral Resources by Zone

Table 18.10 summarizes the Measured and Indicated Mineral Resources collectively by sub-zone.  Of the 32 zones, seven of these zones (EW, FW4, FWA, HW, HW5, HWA, MAIN) account for about 71% of the total tons and 85% of the contained ounces.

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Table 18.10   High grade Measured and Indicated Mineral resources by Zone - 2005

Zone	Tons	Cut Grade (oz Au/ton)	Contained ounces (Cut)	Horizontal Width
EW	20,529	6.37	130,781	7.4
EW1	26,537	2.13	56,602	14.3
FW	18,164	5.31	96,446	5.9
FW2	42,797	1.20	51,512	5.8
FW3	24,600	0.64	15,813	8.4
FW3A	24,192	0.34	8,338	17.8
FW3B	18,700	2.45	45,783	9.6
FW4	222,828	0.84	187,223	9.5
FW4A	31,131	0.76	23,525	15.7
FW4B	6,633	1.31	8,689	6.4
FW4C	30,066	0.96	28,872	6.6
FWA	32,919	3.51	115,513	5.8
FWB	9,586	1.45	13,911	6.0
FWC	762	3.06	2,332	5.7
FWD	542	1.03	557	16.3
FWX	6,146	1.55	9,512	4.1
HW	85,844	2.04	175,545	8.9
HW5	535,279	4.62	2,473,206	20.9
HW5A	525	0.81	423	6.1
HW5C	169	13.44	2,271	12.8
HW6	2,895	3.52	10,204	4.0
HW7	92,604	0.79	72,870	5.9
HWA	114,825	4.93	565,830	11.6
HWB	8,757	1.19	10,442	7.3
HWD	12,558	2.03	25,497	14.9
HWX	4,100	1.21	4,955	6.5
HWX1	12,819	1.56	20,034	7.8
HWY	3,551	1.57	5,585	4.7
MAIN	127,342	2.25	286,973	14.2
MAINA	52,091	1.43	74,535	14.0
MAINB	22,722	3.34	75,947	12.1
MAINX	6,496	3.30	21,453	6.4
Grand Total	1,598,709	2.89	4,621,179	10.6

These zones include most of the highest tonnage zones, which are primarily the highest grade, and have a cut average grade of about 3.5 oz Au/ton.

Inferred Mineral Resources additional to the Resources summarized above amount to 742,000 tons at a grade of 1.37 oz Au/ton.

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18.4.4Mineral Reserve Statement for High Grade Zone

The Mineral Reserve estimate is based on Measured and Indicated Resources for which mining plans have been developed, minus losses due to operational constraints plus mining dilution (see next section for explanation of dilution). Those remaining Mineral Resources not included in Mineral Reserves are tabulated and summarized in Table 18.21.

The Proven designation in the High Grade Zone has been given to 18 ore shoots (see Table 18.23) where continuity has been demonstrated by closely spaced drilling and at least one mining excavation along strike or dip. The polygons in the Proven Mineral Reserves average 575 square feet, which represents a Reserve block approximately 24 feet by 24 feet or drilling at approximately 25-foot centres.

If no mining exists within a zone, the Mineral Reserves are classified as Probable with allowable drillhole spacing, in most cases, up to 50 feet. The same economic criteria apply as for Proven Mineral Reserves.

18.4.5High Grade Zone Mineral Reserves

The Proven and Probable Mineral Reserves in the High Grade Zone are summarized in Table 18.11.

Table 18.11 High Grade Zone Mineral Reserves (as of December 31, 2005)

Category	Tons	Grade	Contained
		(oz Au/ton)	Oz Au
Proven (incl. Broken ore)	677,000	2.27	1,538,000
Probable	1,119,000	2.53	2,831,000
Total	1,796,000	2.43	4,369,000

Dilution

The dilution rate of one ft (30cms) on each contact as predicted in the original Feasibility Study was not achieved and a revision was necessary to 1.5 ft (45cms) in 2002. The dilution of 1.5 ft (45cms) was retained for the 2005 Mineral Reserve estimate, and was applied to each contact. A small additional tonnage is also included to account for flatter veins, where Goldcorp experiences higher than normal dilution.

The significant sources of dilution are attributed to the following:

•unexpected irregularities in complex ore geometry;

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•mining very narrow High Grade veins;

•mining flatter (low angle) “north-south” oriented structures; and

•failure of structure parallel dykes.

Higher internal dilution is being encountered in many zones, owing to the nature of the ore geometry. This was addressed in the 2002 Mineral Reserve estimate by increasing the minimum mining width to 7 ft (2.1m). This higher minimum mining width was retained for the 2005 Mineral Reserve estimate. The 2005 estimate continued to attempt to address higher dilution incurred when mining flatter north-south structures by transferring higher dilution into sub-zones made up of some, or nearly all, of these styles of structures. This final dilution (approximately 128,000 tons) is added after the sub-zones have been brought to the 7-ft (2.1m) minimum width.

Production Reconciliation

On a monthly basis, the total mined (broken) tonnage is calculated from excavated volumes. Total ounces are reported from the mill, adjusted for inventory in ore passes, bins, surface stockpile, and material remaining in the stopes. Ounces are assigned to the various stopes where mining occurred based on information gathered during the month. For every blast in ore, five random muck samples are grabbed, and the face/walls are chip sampled. This assay information is summarized for each blast/stope and used as a guide in assigning ounces back to the stopes.

When a lift of a stope is completed, a solid is constructed of both the ore and the excavation. All assay data (chip samples and grab muck samples), as well as geological mapping, are used to determine the configuration of the ore within the excavation. Volumes of ore and excavation are compared to arrive at a dilution for the entire lift of the stope.

Fill dilution is assumed to minimal. In overhand mining, fill dilution is introduced when the stopes are being cleaned before filling. The cleaning process is laborious, and care is taken to maximize recovery, minimizing fill contamination. In underhand mining, fill dilution is thought to be slightly higher than in overhand methods, but overall, not much difference is observed. Mining recovery is considered to be very close to 100%.

At the end of each month, the Mineral Reserve is “clipped” with the areas mined during the month and results are compared to production. The total production to-date compared to estimated mined out Mineral Reserves using the new capping factors (i.e., Mean + 3 SD for all sub-zones, except for HW5, HW7 and HWA where Mean + 4 SD is still used) and 42.8% estimated dilution of mineralisation compare as shown in Table 18.12. Actual mining results show the dilution to be about 43.5% based on the reconciliation of all mining headings since the beginning of the project in early 2000.

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Table 18.12 Total Production Compared to Estimated Mineral Reserves (to November 30, 2005)

	Tons	Grade	Ounces	Dilution (%)
Reserves	1,352,000	2.29	3,090,000	42.8%
Actual Production *	1,313,000	2.24	2,939,000	43.5%

* Note: Totals reflect Goldcorp estimates of mining during December, rather than actual production.

On a project to-date basis, the variance of the actual tons mined to the Mineral Reserve expected tons is about 3%, while the variance on the ounces mined is less than 5%. The revised 2005 Mineral Reserve estimate (see Table 18.11 previously) compares with estimates for previous years as follows:

1999	1,696,000 tons @ 1.37 oz Au/ton for 2.32 million oz
2000	1,799,000 tons @ 1.68 oz Au/ton for 3.02 million oz
2001	1,850,000 tons @ 2.05 g Au/ton for 3.80 million oz
2002	1,957,000 tons @ 2.35 g Au/ton for 4.59 million oz
2003	1,983,000 tons @ 2.22 g Au/ton for 4.40 million oz
2004	1,994,000 tons @ 2.23 g Au/ton for 4.44 million oz

18.4.6Mineral Reserves By Zone

Table 18.13 reports the Proven and Probable Mineral Reserves by sub-zone and Table 18.14, the general distribution by location. Of the total Mineral Reserves, about 50% of the tonnage and 31% of the ounces are in the vein system currently being mined above 37 Level. At a production rate of approximately 250,000 tons per year or about 700 tpd (2005 average), the current total Mineral Reserve base will last about seven years, assuming all Reserves will be eventually mined. The average grade of all of the above 37 Level Mineral Reserves is 1.50 oz Au/ton (51.4 g/t), compared to currently defined total Mineral Reserves at 2.43 oz Au/ton
(83.3 g/t), which includes the much lower grade undeveloped footwall veins.

The Mineral Reserves below the 37 Level make up 50% of the tons and 69% of the ounces at a grade of 3.38 oz Au/ton (115.9 g/t).

The five largest veins (FW4, HW, HW5, HWA_2 and MAIN_1L) consist of about 1,295,000 tons grading 2.68 oz Au/ton (91.9 g/t) for a total of 3.47 million oz, or approximately 72% of the total tons and 80% of the total contained ounces.

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Table 18.13 High Grade Zone Proven and Probable Mineral Reserves by Zone - 2005

Zone	Tons	Cut Grade	Contained ounces	Dilution
		(oz Au/ton)	(Cut)
EW	35,990	3.63	130,781	80.8
EW1	24,726	1.37	33,811	52.5
FW	20,163	4.38	88,308	82.8
FW2	56,076	0.84	47,188	59.6
FW3	19,836	0.46	9,106	26.3
FW3B	24,677	1.86	45,783	38.7
FW4	158,356	0.63	99,350	37.7
FW4A	35,622	0.63	22,266	16.8
FW4B	5,065	1.27	6,443	50.0
FW4C	20,853	0.71	14,773	53.4
FWA	41,596	2.71	112,683	70.6
FWB	14,564	0.96	13,911	73.9
FWC	1,071	2.18	2,332	51.7
FWD	667	0.84	557	23.1
FWX	6,414	0.97	6,201	180.7
HW	142,121	1.24	175,545	68.4
HW5	693,910	3.54	2,457,641	33.8
HW5A	813	0.52	423	142.7
HW5C	209	10.87	2,271	23.7
HWA_1L	12,915	1.05	13,589	49.7
HWA_2	152,486	3.58	545,815	45.8
HWB	5,269	1.53	8,064	41.2
HWD	14,809	1.72	25,497	20.6
HWX	5,855	0.85	4,955	51.9
HWX1	17,348	1.15	20,034	35.4
HWY	5,793	0.96	5,585	70.1
MAIN_1L	148,313	1.27	189,006	41.9
MAIN_2	28,568	3.43	97,967	31.0
MAINA	60,596	1.22	74,194	18.8
MAINB	27,288	2.73	74,385	45.2
MAINX	7,268	2.82	20,514	49.8
Grand Total	1,789,237	2.43	4,348,978	41.1

Note:  Reserves are in-situ material only and exclude broken ore.

Table 18.14 Tonnage and Grade Distribution of High Grade Zone Mineral Reserves - 2005

Zone	Dilution (%)	Tons	Grade (oz Au/ton)	Contained Ounces	Proportion (%)
Above 37 Level	45.4	903,498	1.50	1,358,428	31
Below 37 Level	36.9	885,739	3.38	2,990,550	69
Total	41.1	1,789,237	2.43	4,348,978	100

Note:  Reserves are in-situ material only and exclude broken ore.

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The average true horizontal width of the High Grade Zone Proven and Probable Mineral Reserves, assuming about 3 ft of dilution (90cms), is almost 14 ft (4m). When the external or mining dilution is added to the minimum horizontal Mineral Resource width of
4 ft (1.2m), the total average dilution for the insitu Mineral Resource is about 41%. The external dilution added to the Mineral Resource estimate represents about 23% of the total Mineral Reserve tons, thereby reducing the Measured and Indicated Resource grade from 2.89 oz Au/t  (99 g/t) to 2.43 oz Au/t (83.3 g/t) in the final High Grade Zone Mineral Reserve estimate (not including broken ore).

As referred to previously under “Zone Interpretation and Modelling” (Section 18.1.3), the Mineral Resource and Reserve estimates include parts of polygons which are below the cutoff grade in order to outline coherent mineable blocks. This material constitutes part of the Mineral Reserve estimate.

18.5      Sulphide Zones Mineral Resource and Reserve Estimate (RLC)

18.5.1   Application of Definitions for the Sulphide Zones

The Mineral Resources and Reserves have been categorized in the standard historical manner for the mine.

Measured Mineral Resource (Proven Reserves) requires at least one or more mine openings to confirm continuity, usually with supporting diamond drillhole information. A Measured Resource is projected halfway to the next data point or a maximum of 25 feet above or below a drift and/or stope, on the basis of chip sampling plus diamond drill results where available.

Indicated Mineral Resources (Probable Reserves) consist of an additional projection of 25 feet beyond the limits of the Measured Resources, but is more commonly based on diamond drilling. An Indicated Resource should show geological continuity and may be based on single blocks surrounded by an additional 25 foot projection in all directions for the Inferred category.

The bulk of the Mineral Resources are however drilled at a regular grid spacing of 25 by 25 feet. Irregularly spaced holes may be grouped and averaged into less regularly shaped blocks where necessary. Complex zones which are highly irregular are estimated by plan outlines and the calculated tonnage per vertical foot method is applied.

Inferred Mineral Resources have been estimated in various parts of the mine based on sparse drilling or projections beyond the Indicated Mineral Reserve limits by an additional 25 feet.

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18.5.2Method of Estimation for Sulphide Zones

The Sulphide Zones have historically been estimated using conventional manual methods and include anticipated mining dilution in the initial calculations based on experience in various parts of the mine. Until recently, only one classification system was used, namely Proven, Probable and Possible, all of which included dilution and related directly to the mine head grade. The Measured and Indicated Mineral Resource categories, as currently introduced by the mine for the Sulphide Zones, have dilution included above the 30 Level and are the same as Proven and Probable Mineral Reserves for that part of the mine.

The mineralized intercept is taken as the full width of the structure regardless of grade, as has been done for the High Grade Zone. When the original estimates were made, the minimum width including dilution is taken as 6 ft (1.8m), which is the same minimum width used previously for the High Grade Zone Mineral Reserves. Although the minimum width in the High Grade Zone has been increased to 7 ft (2.1m) and overall dilution to 1.5 ft (45cms) on either wall, no revision has been made to the Sulphide Zones since previous experience supported the original estimates. However, below the 30 Level, a change has been made and an additional 1.5 ft (45cms) of dilution has been added to each wall, which has resulted in a minimum width of 9 ft (2.7m).

The 2-5-10 cutting factor continues to be used on all Sulphide Zones Mineral Resources with some additional cutting done at the discretion of the geologist based on production experience. The only exception is in the Far East Zone drilled from 16 Level, where a Mean plus 1 SD cut is used. A tonnage factor of 11 cubic feet/ton is applied universally throughout the mine. The assay cutoff grade is 0.15 oz Au/ton (5.1 g/t) in the No. 1 Shaft area and 0.20 oz Au/ton (6.9 g/t) for the No. 2 Shaft area.

All assaying above 30 Level was done at the mine laboratory prior to the mine closure in 1995, with outside check assaying by Chemex. Core from recent drilling has been assayed by Chemex and X-RAL/SGS. Although past production from individual stopes has differed from predicted tons and grade, the overall yearly production has come very close to the predicted Mineral Reserve grades with relatively small pluses and minuses from year to year.

Below the 30 Level, where there has been some additional drilling since 1998 in conjunction with drilling of the High Grade Zone, the Mineral Resources have been re-estimated using the computerized polygonal method. During the last two years, a major revision and reinterpretation of the Sulphide Zones incorporating new drilling information has been carried out. Many of the zones from previous years have either been renamed or re-categorized accordingly.

Additional drilling has been successful in expanding areas of known mineralisation, particularly above 30 Level, and adding Mineral Resources in new areas, such as the Far East Zone (from 16

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Level to a depth of about 3,100 ft / 940m). Continued exploration drilling has also intersected down dip extensions of the main Sulphide Zones at depth. The grade and thickness of the new intersections are fairly consistent with the material extracted historically from the upper levels of the mine.

18.5.3   Statistical Analyses

The Sulphide Zone assay information is continuing to be entered into a computerized database in order to complete polygonal Mineral Resource estimates for all the identified zones. For 2005, new statistics were generated for only those zones where a substantial amount of new drilling or geological modelling has been conducted. WGM decided to retain the statistical tables from 2004 in this report for reference/comparison purposes with the new statistical analysis. Tables 18.15 and 18.16 show the statistical analysis for the two-foot normalized composites for the two years.

Table  18.15 Statistics on Sulphide Zones Two-Foot Normalised Uncut Assays (from 2004)

Variable	All Samples	Oz Au/ton
		<0.20	>0.20 <1.00	>1.00
Mean	0.42	0.10	0.37	3.80
Variance	5.05	0.00	0.03	73.42
Median	0.18	0.11	0.31	1.82
Minimum	0.000	0.000	0.200	1.000
Maximum	226.966	0.199	0.999	226.966
No. of Samples	23,543	13,039	9,122	1,382
% of Samples	100.0	55.4	38.7	5.9

Table  18.16 Statistics on Two-Foot Normalised Uncut Assays on Selected Sulphide Zones for 2005*

Variable	All Samples	Oz Au/ton
		<0.20	>0.20 <1.00	>1.00
Mean	0.43	0.11	0.39	3.25
Variance	2.59	0.00	0.03	34.07
Median	0.21	0.12	0.32	1.68
Minimum	0.00	0.00	0.20	1.00
Maximum	77.10	0.199	0.9998	77.10
No. of Samples	4,428	2,115	2,046	267
% of Samples	100.0	47.8	46.2	6.0

*Includes zones which have received a reasonable amount of recent Mineral Resource modelling work. These zones include the “Deep Sulphide” zones (SC-ESC, ESC-HW and ESC3J), other near-by “ESC3” zones (ESC3B, ESC3C, ESC3D, ESC3E, ESC3F and ESC3G), “Lower E” zones (E-L, E-L-HW and E-L-FW), F Main Zone, and the “Far East” zones (FW-FW1, FE-FW2, FE-FW3, FE-FW4, FE-FW5, FE-HW1 and FE-HW2). The “Upper D and “Upper E” zones and “SC-ESC/ESC-HW Complex” above 28 level intercepts were not included in the 2005 dataset, since they have not received any substantial amount of work over the last several years.

Figures 18.7 and 18.8 graphically illustrate the sample population to be a typical log normal distribution. The mean of all samples is 0.43 oz Au/ton (14.7 g/t) and the maximum is

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77.1 oz Au/ton (2,643.5 g/t). Only about 6% of the composites exceed 1.0 oz Au/ton (34.3 g/t)  as compared to almost 40% for the High Grade Zone.

Tables 18.17, 18.18, 18.19 and 18.20 are presented below for comparison purposes to the High Grade Zone and to illustrate the differences in the gold distribution between the two zones for 2004 and 2005.

Table 18.17  Basic Statistics on Sulphide Zones Raw Assays (from 2004)

Variable	All Samples		Oz Au/ton
		<0.20	>0.20 <1.00	>1.00
Mean	0.68	0.10	0.38	6.02
Variance	35.17	0.00	0.04	411.01
Median	0.18	0.10	0.31	2.02
Minimum	0.001	0.001	0.200	1.000
Maximum	567.000	0.199	0.999	567.000
No. of Samples	18,402	9,802	7,137	1,463
% of Samples	100.0	53.3	38.8	8.0

Table 18.18 Basic Statistics on Raw Assays (Selected Sulphide Zones, 2005 *)

Variable	All Samples		Oz Au/ton
		<0.20	>0.20 <1.00	>1.00
Mean	0.55	0.10	0.39	4.56
Variance	13.69	0.00	0.04	173.83
Median	0.20	0.11	0.33	1.75
Minimum	0.0005	0.0005	0.20	1.00
Maximum	192.72	0.199	1.00	192.72
No. of Samples	3,728	1,818	1,643	267
% of Samples	100.0	48.8	44.1	7.2

*Includes zones which have received a reasonable amount of recent Mineral Resource modeling work. These zones include the “Deep Sulphide” zones (SC-ESC, ESC-HW and ESC3J), other near-by “ESC3” zones (ESC3B, ESC3C, ESC3D, ESC3E, ESC3F and ESC3G), “Lower E” zones (E-L, E-L-HW and E-L-FW), F Main Zone, and the “Far East” zones (FW-FW1, FE-FW2, FE-FW3, FE-FW4, FE-FW5, FE-HW1 and FE-HW2). The “Upper D and “Upper E” zones and “SC-ESC/ESC-HW Complex” above 28 level intercepts were not included in the 2005 dataset, since they have not received any serious amount of work over the last several years.

Table 18.19 Basic Statistics on Sulphide Zones Uncut Mineralised Intercepts (from 2004)

Variable	All Samples		Oz Au/ton
		<0.20	>0.20 <1.00	>1.00
Mean	0.52	0.09	0.39	3.30
Variance	3.99	0.01	0.04	33.74
Median	0.21	0.10	0.32	1.76
Minimum	0.000	0.000	0.200	1.000
Maximum	77.812	0.199	0.997	77.812
No. of Samples	4,303	1,994	1,906	403
% of Samples	100.0	46.3	44.3	9.4

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Table 18.20 Basic Statistics on Uncut Mineralised Intercepts (Selected Sulphide Zones, 2005*)

Variable	All Samples		Oz Au/ton
		<0.20	>0.20 <1.00	>1.00
Mean	0.62	0.15	0.37	3.24
Variance	2.74	0.00	0.02	18.63
Median	0.30	0.16	0.32	1.66
Minimum	0.00	0.00	0.20	1.00
Maximum	21.50	0.199	0.99	21.50
No. of Samples	651	131	452	68
% of Samples	100.0	20.1	69.4	10.4

*Includes zones which have received a reasonable amount of recent Mineral Resource modeling work. These zones include the “Deep Sulphide” zones (SC-ESC, ESC-HW and ESC3J), other near-by “ESC3” zones (ESC3B, ESC3C, ESC3D, ESC3E, ESC3F and ESC3G), “Lower E” zones (E-L, E-L-HW and E-L-FW), F Main Zone, and the “Far East” zones (FW-FW1, FE-FW2, FE-FW3, FE-FW4, FE-FW5, FE-HW1 and FE-HW2). The “Upper D and “Upper E” zones and “SC-ESC/ESC-HW Complex” above 28 level intercepts were not included in the 2005 dataset, since they have not received any serious amount of work over the last several years.

18.5.4   Application of Mineral Reserve Definitions for the High Grade Zone

The Mineral Reserve estimate is based on Measured and Indicated Resources for which mining plans have been developed, minus losses due to operational constraints plus mining dilution (see next section for explanation of dilution). Those remaining Mineral Resources not included in Mineral Reserves are tabulated and summarized in Table 18.21 .

A mining plan is required to show that the zone is economically viable and polygons that do not appear to be economic at present, but which may be shown to be economic by further drilling or mining development, have been kept in the Indicated and Inferred resource categories. The Engineering and Geology departments refine this mining plan each year.

The Proven designation in the High Grade Zone has been given to 18 ore shoots (see Table 18.24) where continuity has been demonstrated by closely spaced drilling and at least one mining excavation along strike or dip. The polygons in the Proven Mineral Reserves average 575 square feet, which represents a Reserve block approximately 24 feet by 24 feet or drilling at approximately 25-foot centres.

If no mining exists within a zone, the Mineral Reserves are classified as Probable with allowable drillhole spacings, in most cases, up to 50 feet. The same economic criteria apply as for Proven Mineral Reserves.

18.5.5   Sulphide Zone Mineral Resources and Reserves

Continuing exploration of the High Grade Zone and Goldcorp’s long-term strategy to support the expanded production rate when the new shaft is operational, has resulted in additions to the

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Mineral Resources and Reserves in the adjacent Sulphide Zones and the Far East Zone (added to the Resources in 2002). These zones have been expanded along strike and at depth below the previously mined level (4,400 ft/ 1,340m).

With the most recent in-fill and deep drilling, the Far East Zone now has a vertical extent stretching from 16 Level to depth of at least 3,600 ft (1,100m). Further drilling from the 26 and 34 Levels is planned to expand and confirm this Mineral Resource.

The Sulphide Zone Mineral Resources and Reserves are summarized and presented by zone in Tables 18.21 to 18.25, respectively. As compared to 2004, the total Measured and Indicated Mineral Resources have decreased by about 35,000 ounces, with a corresponding decrease in tonnage of about 83,000 tons. This is due primarily to a decrease in the Mineral Resources above 30 Level.

Table 18.21 Summary of Mineral Resources of the Sulphide Zones (as of December 31, 2005)

Area	Tons	Grade	Contained ounces
		(oz Au/ton)	of Gold
Surface to 30 Level
Measured	490,000	0.39	190,000
Indicated	1,096,000	0.46	501,000
Sub-total	1,586,000	0.44	691,000
Inferred	272,000	0.51	138,000
Below 30 Level
Measured	0		0
Indicated	1,397,000	0.37	511,000
Sub-total	1,397,000	0.37	511,000
Inferred	743,000	0.55	411,000
Total Measured	490,000	0.39	190,000
Total Indicated	2,493,000	0.41	1,012,000
Total Measured and Indicated	2,983,000	0.40	1,202,000
Total Inferred	1,015,000	0.54	549,000

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Table 18.22 Sulphide Measured and Indicated Mineral Resources by Zone - 2005

Zone	Tons	Cut Grade	Contained ounces	Horizontal
		(oz Au/ton)	(Cut)	Width
D	50,041	0.42	20,799	5.9
D-F	25,933	0.46	12,049	7.6
D-H2	17,819	0.49	8,646	9.6
D-H3	68,457	0.37	25,410	12.4
D-H4	14,541	0.49	7,136	8.0
E-L	133,632	0.52	70,066	15.7
E-L-FW	8,626	0.40	3,445	5.6
E-L-HW	54,294	0.57	30,711	8.7
ESC11	6,133	0.48	2,969	7.2
ESC3B	14,021	0.54	7,585	12.2
ESC3D	24,130	0.35	8,547	15.0
ESC3E	20,464	0.29	5,958	16.1
ESC3J	45,002	0.40	17,976	8.8
ESC4	21,951	0.86	7,795	7.9
ESC9	12,233	0.35	4,221	8.5
ESC-HW	1,010,403	0.41	413,310	14.1
F	9,816	0.49	4,841	8.2
FE-FW2	113,638	0.32	36,818	13.4
FE-FW3	190,249	0.48	91,224	14.2
FE-FW4	7,003	1.00	6,988	5.7
FE-HW2	20,330	0.48	9,817	7.6
FMAIN	37,112	0.73	27,026	6.8
H	69,440	0.33	22,917	17.9
J	7,407	0.48	3,551	16.6
NC	45,086	0.31	14,057	8.7
NS	11,531	0.42	4,873	6.8
PLM	87,306	0.29	25,263	12.6
SC-ESC	629,413	0.34	216,941	10.6
SC-FW2	7,166	1.17	8,369	5.5
SC-HW1	17,802	0.49	8,646	14.1
SC-HW2	154,247	0.32	49,239	11.3
SC-HW3	17,521	0.73	12,720	10.2
SC-HW4	6,183	0.46	2,848	7.2
SS	24,207	0.38	9,095	13.1
Grand Total	2,983,135	0.40	1,201,854	11.7

Of the 34 Sulphide Zones, six (E-L, ESC-HW, FE-FW2, FE-FW3, SC-ESC, and SC-HW2) make up about 73% of the contained ounces and 75% of the tons. These include most of the highest tonnage zones, which are not necessarily the highest grade, and have a cut average of about 0.39 oz Au/ton (13.4 g/t).

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Table 18.23 Summary of Mineral Reserves of the Sulphide Zones (as of December 31, 2005)

Area	Tons	Grade	Contained ounces
		(oz Au/ton)	of Gold
Surface to 30 Level
Proven	204,000	0.40	82,000
Probable	416,000	0.46	193,000
Subtotal	620,000	0.44	275,000
Below 30 Level
Proven	0		0
Probable	706,000	0.39	272,000
Subtotal	706,000	0.39	272,000
Total Proven	204,000	0.40	82,000
Total Probable	1,122,000	0.41	465,000
Total Sulphide Reserves	1,326,000	0.41	547,000

Table 18.24 Sulphide Zones Proven and Probable Mineral Reserves by Zone Above Level 30 (as of December 31, 2005)

Zone	Tons	Cut Grade	Contained ounces	Average Width	Dilution
		(oz Au/ton)	(Cut)	(feet)	(%)
D-F	31,143	0.33	10,146	11.5	43.1
D-H2	23,506	0.37	8,646	12.7	32.2
D-H3	19,027	0.44	8,313	12.3	34.6
E-L	137,742	0.47	65,112	18.8	24.4
E-L-HW	61,443	0.42	26,103	13.2	35.4
ESC3B	470	0.23	110	13.1	30.6
ESC-HW	93,728	0.49	46,364	13.4	35.3
F	7,954	0.40	3,191	11.7	34.3
FMAIN	53,826	0.50	27,026	9.9	64.1
H	17,145	0.35	6,039	41.0	7.9
SC-ESC	92,457	0.43	40,208	13.5	36.0
SC-FW2	11,160	0.75	8,369	8.6	64.7
SC-HW1	8,665	0.55	4,800	14.5	26.9
SC-HW2	13,666	0.31	4,258	13.6	30.3
SC-HW3	9,458	0.48	4,556	14.9	26.8
SC-HW4	8,822	0.32	2,848	10.3	78.6
SS	29,758	0.31	9,095	16.1	30.4
Grand Total	619,970	0.44	275,183	13.9	34.4

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Table 18.25 Sulphide Zones Proven and Probable Mineral Reserves by Zone Below Level 30 (as of December 31, 2005)

Zone	Tons	Grade Cut (oz Au/ton)	Contained ounces (Cut)	Average Width (feet)	Dilution (%)
ESC3B	9,512	0.61	5,787	13.0	31.0
ESC3D	29,071	0.29	8,547	18.1	20.5
ESC3J	60,751	0.30	17,976	11.9	38.3
ESC-HW	497,475	0.42	209,407	16.1	23.6
SC-ESC	109,373	0.28	30,386	15.1	27.0
Grand Total	706,182	0.39	272,103	15.5	25.3

Inferred Mineral Resources additional to the Resources summarized above amount to 1,015,000 tons at a grade of 0.54 oz Au/ton (18.5 g/t).

18.6Campbell Complex

The Mineral Resources and Mineral Reserves for the Campbell Complex are reported as of September 30, 2006. The Mineral Resource Statement was calculated with the supervision of WGM. Currently, the Campbell Complex is revising the methodology for resource estimation. Previously, under Placer Dome, Campbell mine resources were derived from block models utilizing the in-situ  width of mineralization and a global cutoff to direct resource interpretation and estimation. A transition is ongoing to dilute all block models to a minimum 6 ft (1.8m) mining width, develop project specific cutoff grades such that the remaining resource (after removal of reserves) is more reflective of the resource grade located inside the reserves, and a revised resource classification is being utilized to produce more continuous areas of similar classed resources while decreasing local variability in resource class.

Computerized Campbell Complex resources have been calculated using two methods. The “DIGS computerized method” (referred to as DIGS blocks) was the first computerized system for reserve/resource estimation, is deemed historical, and has since been replaced by block modeling. Datamine® software is currently used for grade estimation and was predated by Vulcan®.

18.6.1Mineral Resource Statement (CC)

The Mineral Resource Statement dated September 30, 2006 is comprised of 63 projects, which commonly include multiple zones. The Mineral Resource Statement reflects only the remaining resources exclusive of mineral reserves.

Please note that mineral resources that are not mineral reserves do not have demonstrated economic viability.

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Two Mineral Resource Statements are listed below, the first reflecting the current Goldcorp methodology (block models from Datamine®), and the second reflecting historical resources (dated to the year 2000) using the old DIGS resource method.

Table 18.26 Mineral Resource Statement for Campbell Complex to September 2006*

	Short Tons	oz/ton	Ounces
Measured	229,731	1.13	258,535
Indicated	616,596	0.55	338,932
Sub Total	846,327	0.71	597,466
Inferred	1,310,797	0.87	1,134,216

*derived from Datamine block models exclusive of reserves.

Table 18.27 Digs Resource Statement for residual material*

Class	Short Tons	Oz/ton	Ounces
Measured	131,644	0.75	98,675
Indicated	96,464	0.63	60,687
Sub Total	228,107	0.70	159,362
Inferred	141,571	0.64	90,545

*Cautionary statement: DIGS resources are considered historical and should not be relied upon.

The resources reported in the DIGS Resource Statement is the residual amount of material from the figure last calculated in the year 2000. These figures have been re-calculated for Table 18.27 to allow a minimum 6 ft (1.8m) mining width and reported with a cutoff of 0.35 oz/ton (12 g/t).

18.6.2Mineral Reserve Statement (CC)

Table 18.28 Mineral Reserve Statement for Campbell Complex

Campbell complex Reserves as of Sept. 31-06

	Prov. Prob. Tonnes (000’s)	Grade (g/t)	Prov. Prob. KGrams
Reserves UG	1,985	16.25	32,256
Reserves Tailings	2,931	1.73	5,082
Total Reserves	4,917	7.59	37,337

	Prov. Prob. Tons (000’s)	Grade (oz/t)	Oz
Reserves UG	2,189	0.47	1,037,046
Reserves Tailings	3,231	0.05	163,380
Total Reserves	5,420	0.22	1,200,426

- 89 - November 17, 2006

18.6.3Mineral Resource Methodology

Block Model Resources

As the transition to diluted models is currently in progress, the initial five projects for remodelling were selected as they represent the largest ounces contributions to the mineral resources (including resources in reserves) as of December 31, 2005.

Following creation of the diluted model, the blocks are flagged as either unmined, mined or not in resource (unavailable). The unmined blocks which fall within the mine plan are then flagged as reserves. The resources flagged as reserves are tabulated to determine an average reserve grade, with a target grade of 75% of the reserve grade established for the remaining resources.

The newly adopted method for determining remaining resources is to set a target grade at approximately 75% of the reserve grade for a given project, to account for positive items such as synergies from integration of the Campbell and Red Lake Complexes, improved mining techniques, and possible upside in the gold price. A project specific cutoff grade is then applied to achieve the target grade, and determine the tons and grade of the remaining resource above cutoff. For the five models, factors reflecting the percentage of change to the resource tons and ounces were derived. These factors have been globally applied on the remaining projects to estimate the remaining resource tons and ounces. This methodology is used to produce a resource statement of diluted tons and grade comparable to the reserves statement.

Cap grades for the five major projects (by measured and indicated ounces) ranged from 0.21 oz/ton to 0.38 oz/ton (7.2 g/t to 13 g/t) Au, with an unweighted average cutoff grade of 0.27 oz/ton (9.3 g/t) gold.

DIGS Resource Method

The DIGS resource method (DIGS blocks) was used up to, and including the year, 2000 and is deemed as historical. The DIGS method was the first computer based resource/reserve methodology to be used by Placer Dome at Campbell mine, which was succeeded by block modeling in 2001. DIGS resource blocks are continually being removed as they are replaced by new projects or extensions of existing project areas encompassed by block models.

The DIGS resources were created using a polygonal based methodology. Based on geological mapping from levels and sublevels, and available drill and chip data, top and bottom “chip strings” were constructed. Two-dimensional planes representing the auriferous structures, referred to as “DIGS blocks”, were then built from the chip strings extending the known

- 90 - November 17, 2006

mineralization away from development. Samples were manually flagged and composited referencing the appropriate ore block.

Tonnages were calculated based the average of the top and bottom chip string lengths, multiplied by average composite length, and the tonnage factor applied. The average of the capped composite grades was applied to the block. Cap grades were set using the 10-5-2 method (i.e. a 12 oz/ton grade would be capped to 10 oz/ton).

18.6.4     Mineralized Zone Model (CC)

Resource estimates from 63 block models are incorporated into the September 30, 2006 Mineral Resource Statement. Due to proximity of the zones it is common to have more than one ore zone with in a project. The ore solids (wireframes) representing the mineralisation envelopes (ore structures) were constructed in a three dimensional environment utilizing both plan and section views during the creation process. The building and naming of ore solids are influenced by geology interpretations; lithological units, structures, faults and mineralization. Ore solids are constructed as undiluted in-situ solids.

Diluted block models were created on the five projects with the largest ounce contributions (pre-reserves) at December 31, 2005. Table 18.28 illustrates the ounce contributions of the top ten of the remaining 53 projects.

Table 18.29   Block Model Remaining Resources (Excluding Reserve)

		Measured + Indicated Resource			Oz.% of total
	Project	Tons	Grade	Oz	remaining
	36_30_MM_TP	73,229	0.431	31,542	5.3%
	10_00_F	92,043	1.045	96,200	16.1%
Diluted models	42_36_DC	77,100	0.675	52,017	8.7%
	308_27_56_561	58,031	0.990	57,458	9.6%
	42_36_MM_TP_SXC	120,892	0.315	38,117	6.4%
Remaining models	25_21_G_61_62	61,699	0.461	28,472	4.8%
	58_39_TP_PCB	57,088	0.641	36,620	6.1%
	142_10_F	3,235	3.223	10,427	1.7%
	16_13_L	17,562	0.751	13,196	2.2%
	27_22_56_561_E	7,344	2.510	18,434	3.1%
	Remaining 53 Projects	278,117	0.773	214,985	36.0%
Total Remaining Resource		846,340	0.706	597,469	100%

- 91 - November 17, 2006

18.6.5     Compositing and Capping

Following solid construction an automated data selection process is run which selects drill and chip intercepts encompassed in the solids. Visual examination of both drill and chip data against the wireframes is then undertaken to ensure correct samples are used.

Compositing at the Campbell Complex is reviewed on a project by project basis, with the composite length variable within each project, by zone and sample type (chip or drill hole). The composite interval is chosen to be approximately equal to the mean length or sample interval representing the largest cumulative length of the sampling. Tables 18.30 and 18.31 show an analysis of the univariate statistics for the uncapped composites for both the chip data and the drillhole data.

Table 18.30   Statistics for Five Major Projects - Uncapped Composite Chip Data

		Chip Univariate Statistics
Project	Zone	Mean	Median	Min	Max	Count	Std Dev	Coef of Var
10_00_F	F	1.12	0.14	0.00	306.54	6116	5.52	4.91
	84	0.22	0.02	0.00	9.83	503	1.00	4.51
	K	1.03	0.05	0.00	44.98	478	3.82	3.63
308_27_56_561	56	0.31	0.03	0.00	21.00	721	1.66	5.30
	561	2.54	0.04	0.00	398.19	1194	15.32	6.04
	561HW	0.17	0.00	0.00	1.75	12	0.48	2.90
	56HW	-	-	-	-	-	-	-
36_30_mm_tp	TP	0.81	0.11	0.00	56.73	579	4.76	5.86
42_36_MM_TP_SXC	MM	0.11	0.00	0.00	2.00	18	0.46	4.12
	TP	0.31	0.14	0.00	5.15	114	0.62	2.01
	SXC	-	-	-	-	-	-	-
42_36_DC	DC	1.97	0.40	0.00	276.20	1454	8.50	4.31

- 92 - November 17, 2006

Table 18.31   Statistics for Five Major Projects - Uncapped Composite Drillhole Data

		Drill Hole Uninvariate Statistics
Project	Zone	Mean	Median	Min	Max	Count	Std Dev	Coef of Var
10_00_F	F	0.38	0.00	0.00	50.00	950	3.04	7.98
	84	0.11	0.00	0.00	6.01	178	0.52	4.78
	K	0.55	0.00	0.00	57.24	249	4.41	8.08
308_27_56_561	56	0.07	0.00	0.00	7.03	721	0.33	4.75
	561	0.60	0.00	0.00	126.69	1308	4.26	7.15
	561HW	0.15	0.00	0.00	5.66	122	0.61	4.01
	56HW	0.48	0.00	0.00	30.75	84	3.37	7.09
36_30_mm_tp	TP	0.25	0.10	0.00	26.09	2847	0.96	3.80
42_36_MM_TP_SXC	MM	0.23	0.09	0.00	11.92	123	1.09	4.68
	TP	0.26	0.12	0.00	46.00	1861	1.34	5.12
	SXC	0.30	0.14	0.00	7.52	332	0.70	2.36
42_36_DC	DC	0.77	0.06	0.00	56.65	1367	2.98	3.89

Similarly to the composite intervals, cap grades are also determined on a project, zone, and sample type basis. Selection of cap grades is guided by univariate statistics, grade histograms, probability plots, and decile analysis. Cap grades from projects with extensions of the same zone may be examined, and geological input is sought from appropriate personnel to select the cap grade. Figures 18.9 through 18.28 are a series of probability plots and histograms showing how the cap grade was determined for each of the projects reviewed. Table 18.32 shows the capping levels for composited assays from the 5 major projects that were analysed.

Table 18.32   Capping Levels for Composited Assays for the Five Major Projects

Project	Zone	Sample Type	Dec.31, 2005 cap grades (oz/ton Au)	Current cap grades (oz/ton Au)	Percentage Capped Composites	Percentage Metal Loss
10_00_F	F	hole	38	45	0.09	0.99
	F	chip	38	38	0.22	9.33
	K	hole	38	3.5	0.92	69.71
	K	chip	38	6	1.88	20.36
	84	hole	38	4.5	1.12	19.09
	84	chip	38	13	0.94	14.4

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Table 18.32   Capping Levels for Composited Assays for the Five Major Projects

Project	Zone	Sample Type	Dec.31, 2005 cap grades (oz/ton Au)	Current cap grades (oz/ton Au)	Percentage Capped Composites	Percentage Metal Loss
308_27_56_561	56	hole	40	2.5	0.28	9.78
	56	chip	40	15	0.14	2.74
	56HW	hole	40	2.5	2.38	76.9
	56HW	chip	40	no cap*	-	-
	561	hole	40	17	0.79	21.44
	561	chip	40	60	1.05	22.55
	561HW	hole	40	2.5	0.82	14.24
	561HW	chip	40	2.5	0	0
36_30_mm_tp	TP	hole	7.5	13	0.17	3.22
	TP	chip	7.5	11	1.02	40.25
42_36_mm_tp_sxc	TP	hole	4.5	4	0.64	16.29
	TP	chip	4.5	2	2.63	10.66
	MM	hole	1.2	1	1.85	23.87
	MM	chip	1.2	1	5.88	36.27
	SXC	hole	0.6	2.5	1.5	11.45
	SXC	chip	0.6	no cap*	-	-
42_36_DC	DC	hole	29	40	0.22	1.89
	DC	chip	29	35	0.34	10.73

* no composites

18.6.6     Grade Estimation Methods

During the initial block model creation, a rotation is applied to align the model with the orientation of the mineralized structures. Block size is influenced by active mining practices, and sub-blocking permitted for optimal block filling of thewireframes.

Due to zone variability of the mineralizedstructures, search parameters, and number of composites (minimum and maximum) in the estimation process are customized on a project and zone basis. Ellipsoidal search parameters are determined based on geological evidence from historical mining, recent development within the project area and geological confidence. Both

- 94 - November 17, 2006

chip and hole composites are used in the current estimation process which utilizes inverse distance weighting.

Resource classification utilizes ellipsoidal search, on established search radii, octant search, and minimum sample requirements. Resource classification parameters are established during the estimation process, and are adjusted on a project basis reflecting geological continuity and confidence. See project parameters below for the five major projects..

As part of the approval process block models and results are visually validated and discussed with appropriate geological personnel. Included is the querying of composite intervals against block grades. Resource classification parameters and cap grades may be adjusted reflecting the confidence or knowledge of geological personnel familiar with the project area at which time the model would be re-estimated and subsequently re-examined.

18.6.7     Verification of resource estimation procedures - Model: 10_00_F (includes 3 zones)

Parameter	Value
Estimation method	Inverse Distance
Search ellipse	F Zone :Ellipse, 40 ft (X), 128 ft (Y), 256 ft (Z), rotation: -55° (Z), -5° (X), -10° (Y)
	Octant search
	K Zone :Ellipse, 40 ft (X), 128 ft (Y), 256 ft (Z), rotation: 30° (Z), 20° (X), -20° (Y)
	Octant search
	84 Zone :Ellipse, 40 ft (X), 128 ft (Y), 256 ft (Z), rotation: -85° (Z), -5° (X), -10° (Y)
	Octant search
Restrictions	F, K, and 84 Zones: Minimum number of composites: 3
	Maximum number of composites: 24
	Minimum octants 2, minimum 1 and maximum 4 per octant
	Ellipsoid expansion: up to 2 times, doubling then tripling the ellipsoid radii.

F and K Zones

Measured resources: block that has at least 3 composites, and 2 octants within a search ellipsoid 10 ft by 16 ft by 32 ft oriented like the interpolation search ellipsoid.

Indicated resources:  block that has at least 3 composites, and 2 octants within a search ellipsoid 40 ft by 40 ft by 80 ft oriented like the interpolation search ellipsoid.

Inferred resources: block that has at least 3 composites, and 2 octants within a search ellipsoid 40 ft by 128 ft by 256 ft oriented like the interpolation search ellipsoid.

- 95 - November 17, 2006

84 Zone

Measured resources: block that has at least 3 composites, and 2 octants within a search ellipsoid 10 ft by 16 ft by 32 ft oriented like the interpolation search ellipsoid.

Indicated resources:  block that has at least 3 composites, and 2 octants within a search ellipsoid 10 ft by 40 ft by 64 ft oriented like the interpolation search ellipsoid.

Inferred resources: block that has at least 3 composites, and 2 octants within a search ellipsoid 40 ft by 128 ft by 256 ft oriented like the interpolation search ellipsoid.

18.6.8     Verification of resource estimation procedures - Model: 308_27_56_561 (includes 4 zones)

Parameter	Value
Estimation method	Inverse Distance
Search ellipse	56 and 56HW Zones :Ellipse, 40 ft (X), 128 ft (Y), 256 ft (Z), rotation: -45° (Z), -30° (X), -30° (Y)
	Octant search
	561 and 561HW Zones :Ellipse, 40 ft (X), 128 ft (Y), 256 ft (Z), rotation: -35° (Z), -10° (X), -27°
	(Y), Octant search
Restrictions	56 and 56HW Zones: Minimum number of composites: 1
	Maximum number of composites: 20
	Minimum octants 1, minimum and maximum 3 per octant
	Ellipsoid expansion: up to 2 times, doubling then tripling the ellipsoid radii.
	561 and 561HW Zones: Minimum number of composites: 1
	Maximum number of composites: 20
	Minimum octants 1, minimum 1 and maximum 1 per octant
	Ellipsoid expansion: up to 2 times, doubling then tripling the ellipsoid radii.

56 and 56HW Zones

Measured resources: block that has at least 10 composites, minimum 1 per octant and maximum 3 per octant within a search ellipsoid 5 ft by 16 ft by 32 ft oriented like the interpolation search ellipsoid.

Indicated resources:  block that has at least 4 composites, minimum1 per octant and maximum 3 per octant within a search ellipsoid 10 ft by 50 ft by 64 ft oriented like the interpolation search ellipsoid.

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Inferred resources: block that has at least 1 composite, and minimum 1 per octant within a search ellipsoid 40 ft by 128 ft by 256 ft oriented like the interpolation search ellipsoid.

561 and 561HW Zones

Measured resources: block that has at least 10 composites, and minimum1 per octant within a search ellipsoid 5 ft by 16 ft by 32 ft oriented like the interpolation search ellipsoid.

Indicated resources:  block that has at least 3 composites, and minimum 1 per octant and maximum 1 per octant within a search ellipsoid 10 ft by 50 ft by 64 ft oriented like the interpolation search ellipsoid.

Inferred resources: block that has at least 1 composite, and minimum 1 per octant within a search ellipsoid 40 ft by 128 ft by 256 ft oriented like the interpolation search ellipsoid.

18.6.9     Verification of resource estimation procedures - Model: 36_30_mm_tp

Parameter	Value
Estimation method	Inverse Distance
Search ellipse	TP Zone:Ellipse, 20 ft (X), 128 ft (Y), 256 ft (Z), rotation: -45° (Z), -20° (X), -30° (Y)
	Octant search
Restrictions	TP Zone: Minimum number of composites: 1
	Maximum number of composites: 20
	Minimum octants 1, minimum 1and maximum 4 per octant
	Ellipsoid expansion: up to 2 times, doubling then tripling the ellipsoid radii.

Measured resources: block that has at least 9 composites, minimum 1 per octant and maximum 4 per octant within a search ellipsoid 20 ft by 32 ft by 32 ft oriented like the interpolation search ellipsoid.

Indicated resources:  block that has at least 7 composites, minimum1 per octant and maximum 4 per octant within a search ellipsoid 20 ft by 70 ft by 70 ft oriented like the interpolation search ellipsoid.

Inferred resources: block that has at least 3 composites, and minimum 1 per octant within a search ellipsoid 40 ft by 128 ft by 256 ft oriented like the interpolation search ellipsoid.

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18.6.10  Verification of resource estimation procedures - Model: 42_36_DC

Parameter	Value
Estimation method	Inverse Distance
Search ellipse	DC Zone:Ellipse, 30.6 ft (X), 82.3 ft (Y), 126.1 ft (Z), rotation: -45° (Z), -35° (X), 180° (Y)
	Octant search
Restrictions	DC Zone: Minimum number of composites: 1
	Maximum number of composites: 20
	Minimum octants 1, minimum 1and maximum 3 per octant
	Ellipsoid expansion: up to 2 times, doubling then tripling the ellipsoid radii.

Measured resources: block that has at least 10 composites, minimum 1 per octant and maximum 3 per octant within a search ellipsoid 20 ft by 32 ft by 32 ft oriented like the interpolation search ellipsoid.

Indicated resources:  block that has at least 7 composites, minimum1 per octant and maximum 3 per octant within a search ellipsoid 20 ft by 80 ft by 80 ft oriented like the interpolation search ellipsoid.

Inferred resources: block that has at least 1 composite, and minimum 1 per octant and maximum 3 per octant within a search ellipsoid 40 ft by 128 ft by 256 ft oriented like the interpolation search ellipsoid.

18.6.11  Verification of resource estimation procedures - Model: 42_36_mm_tp_sxc (includes 3 zones)

Parameter	Value
Estimation method	Inverse Distance
Search ellipse	TP Zones: Ellipse, 20 ft (X), 64 ft (Y), 128 ft (Z), rotation: -40° (Z), -20° (X), -30° (Y) Octant
	search
	MM Zone: Ellipse, 20 ft (X), 64 ft (Y), 128 ft (Z), rotation: -25° (Z), -30° (X), -50° (Y) Octant
	search
	SXC Zone: Ellipse, 20 ft (X), 64 ft (Y), 128 ft (Z), rotation: -20° (Z), -20° (X), -30° (Y) Octant
	search
Restrictions	TP, MM and SXC Zones: Minimum number of composites: 1
	Maximum number of composites: 20
	Minimum octants 1, minimum 1and maximum 4 per octant
	Ellipsoid expansion: up to 2 times, doubling then tripling the ellipsoid radii.

- 98 - November 17, 2006

TP, MM, and SXC Zones

Measured resources: block that has at least 10 composites, minimum 4 per octant and within a search ellipsoid .01 ft by .01 ft by .01 ft oriented like the interpolation search ellipsoid.

Indicated resources:  block that has at least 7 composites, minimum 3 per octant within a search ellipsoid 20 ft by 80 ft by 80 ft oriented like the interpolation search ellipsoid.

Inferred resources: block that has at least 3 composites, and minimum 1 per octant within a search ellipsoid 40 ft by 128 ft by 256 ft oriented like the interpolation search ellipsoid.

- 99 - November 17, 2006

Table 18.33   Project 10_00_F - F Zone Significant Drill Intercepts

					True
			Gold Grade	Length	Thickness
Hole Number	From (feet)	To      (feet)	(oz/ton)	(feet)	(feet)
D06240	156	164.6	23.69	8.6	6.79
D07353	42	46	18.01	4	3.69
D04258	80	92.5	13.30	12.5	11.59
D07380	39	48.7	11.03	9.7	8.35
D093R	10.5	14.5	9.06	4	3.40
D07487	49	54.5	5.83	5.5	3.16
D07430	68.6	82	5.32	13.4	8.55
D06605	20	28.5	4.20	8.5	7.10
D08341	89.2	96.5	3.76	7.3	2.12
D08390	0	1	2.46	1	0.82
D08234	10.1	14.8	2.45	4.7	4.41
D09264	65	70	1.99	5	4.22
D09236	129	130	1.78	1	0.90
D05178	28	44.3	1.67	16.3	12.07
D06233	141.5	149	1.43	7.5	6.85
D06385	151.5	157	1.32	5.5	3.80
D09236	24.5	30.3	1.32	5.8	5.23
D07614	0	4	1.20	4	3.85
D03313	12	14.2	1.17	2.2	1.71
D05207	0	5.5	1.10	5.5	1.22
D08388	0	0.5	1.08	0.5	0.16
D095R	22.5	29.5	1.08	7	3.21
D05272	27.5	30.6	1.07	3.1	2.92
D05389	0	1	1.04	1	0.90
D07432	88	91.5	0.99	3.5	2.77
D08335	0	7.5	0.96	7.5	6.11
D08349	22.5	34.5	0.92	12	9.18
D04271	9.7	20.5	0.86	10.8	10.08
D06490	0	7.1	0.76	7.1	6.37
D0824A	14.3	19.1	0.74	4.8	2.96
D05179	28.5	40	0.73	11.5	8.69
D08191	36.1	42	0.72	5.9	5.48
D07509	79.3	85.4	0.70	6.1	4.98

- 100 - November 17, 2006

					True
			Gold Grade	Length	Thickness
Hole Number	From (feet)	To      (feet)	(oz/ton)	(feet)	(feet)
D07320	28.9	37.5	0.66	8.6	7.85
D09309	40	52	0.64	12	11.09
D05273	52.5	62.5	0.64	10	6.54
D096A	28	32	0.58	4	3.65
D07522	142	158.5	0.58	16.5	9.57
D08249	92.5	98.5	0.56	6	2.43
D08297	0	5.9	0.55	5.9	5.43
D05275	14.5	19.5	0.51	5	4.42
D07307	42.5	49.3	0.51	6.8	6.17
D09270	66	66.5	0.50	0.5	0.47
D03265	0	4.2	0.50	4.2	3.73
D08409	0	5.6	0.47	5.6	5.28
D07313	32.3	42.3	0.45	10	9.15
D07536	76.5	83	0.44	6.5	4.78
D08256	2	15.2	0.40	13.2	12.48
D09190	30	33.7	0.39	3.7	2.31
D06606	23	27.8	0.38	4.8	3.80
D09200	4.5	14.5	0.38	10	9.29
D03210	12	17	0.37	5	3.46
D04263	34.5	57.5	0.33	23	16.15
D08259	10.5	24.1	0.32	13.6	5.82
D07535	105	114.5	0.32	9.5	4.88
D08221	33	37.5	0.31	4.5	4.14
D07306	75	80.5	0.31	5.5	5.02
D09205	12	62	0.30	50	12.65
D07647	81.7	90.5	0.28	8.8	0.09
D089A	29.5	35.8	0.27	6.3	4.28
D07351	112.1	128	0.26	15.9	7.16
D09238	18.5	19.5	0.26	1	0.95
D04255	107	130	0.25	23	16.77
D04401	0	3.4	0.25	3.4	2.88

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Table 18.34   Project 36_30_mm_tp - TP Zone Significant Drill Intercepts

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D30155	321.3	322.8	1.5	18.28	1.46
D30335	442	444.6	2.6	13.80	2.43
D30257	150.5	157.6	7.1	10.02	3.00
D33216	688.5	694	5.5	6.03	5.11
D27322	294	300.3	6.3	4.25	5.47
D33216	727	736	9	3.86	8.35
D30335	421.6	423.9	2.3	3.38	2.14
D30145	31.5	34.1	2.6	2.92	2.13
D30334	342.9	351.2	8.3	2.81	7.49
D30392	490.5	493.5	3	2.76	2.04
D27247	1537.6	1542.4	4.8	2.61	4.78
D27138	78.8	82	3.2	2.40	2.82
D30296	0	1.3	1.3	2.22	1.18
D30425	101.6	111	9.4	2.15	5.04
D30342	90.7	94.6	3.9	2.12	3.50
D30223	103.8	108.5	4.7	1.89	2.93
D30386	372.2	377.2	5	1.83	4.24
D30221	63.4	72.1	8.7	1.64	7.95
D30383	435.3	440.8	5.5	1.61	3.25
D30445	5	26.1	21.1	1.49	21.09
D33004	415	418.9	3.9	1.48	3.25
D30226	68.3	71	2.7	1.44	2.18
D30415	78.4	83.6	5.2	1.44	3.41
D27268	64.1	69.5	5.4	1.42	3.36
D33003	894.6	898.3	3.7	1.39	3.28
D27273	255.3	260.9	5.6	1.37	4.65
D27901	55.3	58.6	3.3	1.30	2.98
D30228	69.8	71.4	1.6	1.27	1.46
D30384	476.1	485	8.9	1.14	5.84
D30267	207.4	211.5	4.1	1.13	1.02
D30145	8	21	13	1.10	10.64
D30636	85.3	96.4	11.1	1.10	5.69
D30160	553.8	560.6	6.8	1.09	3.75
D30395	515	525	10	1.05	5.29

- 102 - November 17, 2006

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D30335	345.4	347.6	2.2	1.02	2.04
D30423	245.2	248.7	3.5	1.01	2.76
D30423	64.6	79.2	14.6	0.98	11.32
D33178	444	450	6	0.96	5.87
D30340	258.3	262	3.7	0.94	3.42
D33217	716.5	721.9	5.4	0.92	5.14
D30154	39.9	42.4	2.5	0.84	2.47
D33001	92	97.7	5.7	0.84	4.53
D30817	34	49.5	15.5	0.80	1.78
D30224	71	72.7	1.7	0.80	1.52
D33210	705	709	4	0.80	3.80
D30357	279.8	283.6	3.8	0.76	3.68
D30815	65.5	69.5	4	0.74	0.29
D27266	465	490	25	0.73	21.20
D33210	674	678	4	0.70	3.80
D30423	0	19.1	19.1	0.70	14.72
D30391	419	422.3	3.3	0.66	2.60
D27217	14	16.6	2.6	0.66	2.06
D33217	591.2	597.3	6.1	0.66	5.81
D30431	201	213	12	0.64	4.88
D33144	678	682	4	0.62	3.95
D33215	726.6	730.5	3.9	0.61	3.56
D30154	397	399.8	2.8	0.60	2.77
D27304	100.3	104.7	4.4	0.60	3.92
D27265	310.3	322.8	12.5	0.60	11.65
D30447	11.4	23.2	11.8	0.60	7.92
D30428	109.6	115.6	6	0.59	5.46
D30355	375.1	377.7	2.6	0.57	2.26
D33003	917.2	924.9	7.7	0.57	6.82
D30360	493.5	504	10.5	0.57	7.29
D30448	152.6	156.3	3.7	0.56	1.91
D30420	0	42	42	0.56	32.05
D27300	142.8	148.6	5.8	0.55	2.01
D30419	8	28	20	0.54	18.51
D30421	152.6	159	6.4	0.53	4.01

- 103 - November 17, 2006

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D27376	854.3	858.1	3.8	0.53	3.57
D30229	135.3	141	5.7	0.53	3.42
D33002	136	145.9	9.9	0.52	6.10
D30158	211	216	5	0.52	4.68
D30444	0	42.6	42.6	0.52	16.29
D30450	192.9	198	5.1	0.51	3.66

Table 18.35 Project 42_36_mm_tp_sxc - TP Zone Significant Drill Intercepts

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D36104	212.4	214.2	27.54	1.8	1.72
D39181	338.6	343	21.10	4.4	4.39
D39024	731.7	735	4.80	3.3	2.72
D39227	218	221.5	3.87	3.5	3.48
D39325	78	80	3.70	2	1.43
D36021	144.4	148.1	3.47	3.7	3.06
D39299	268.2	270.8	3.33	2.6	2.26
D36034	579.4	591.5	3.16	12.1	8.42
D36040	65	65.5	2.37	0.5	0.46
D36055	145	146.5	2.24	1.5	1.17
D39025	578	584.8	2.23	6.8	6.24
D36081	505	506.7	2.17	1.7	1.40
D36038	582	584	2.12	2	1.69
D36014	501.6	507.5	2.03	5.9	2.90
D39177	381	384.7	1.95	3.7	3.68
D36057	345.5	352.8	1.72	7.3	2.86
D36062	195.5	197.5	1.43	2	1.62
D39032	869.99	872.49	1.41	2.5	1.42
D36055	204.8	206.8	1.40	2	1.58
D39012	575.7	586	1.37	10.3	8.39
D39326	86.2	88.5	1.21	2.3	1.46
D36053	168.6	172.4	1.19	3.8	3.72
D39022	800.3	802	1.19	1.7	1.02

- 104 - November 17, 2006

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D39305	153	157	1.18	4	3.17
D36058	124.9	128.9	1.13	4	3.75
D39595	0	2	0.99	2	1.97
D39181	376	378.4	0.97	2.4	2.39
D39145	813.1	815.4	0.93	2.3	1.54
D39135	368.9	374.5	0.92	5.6	5.33
D36014	2583.7	2587.3	0.92	3.6	2.25
D39031	770.6	776.7	0.90	6.1	4.00
D39591	35.2	38.5	0.88	3.3	2.93
D39204	468.4	469.8	0.87	1.4	1.14
D39215	233.7	240	0.86	6.3	5.41
D39224	378.2	383.7	0.82	5.5	2.65
D39180	384	385.8	0.80	1.8	1.78
D39049	1282.3	1287.8	0.74	5.5	2.25
D39292	203.6	205.3	0.70	1.7	1.69
D36038	180	182.5	0.70	2.5	2.03
D36039	463	467.5	0.65	4.5	3.23
D39022	865.5	867	0.64	1.5	0.91
D39157	89.3	91.9	0.63	2.6	2.30
D39012	402.3	405.8	0.60	3.5	2.80
D39593	77.2	80.6	0.56	3.4	3.19
D39224	346	355.5	0.55	9.5	4.58
D36052	142.4	150.1	0.53	7.7	7.17
D39549	16.5	18.5	0.52	2	1.99
D39156	163.9	168.2	0.51	4.3	3.79
D39598	190	196.5	0.50	6.5	4.40
D40008	1213	1228.3	0.50	15.3	11.32
D39552	78.5	81.7	0.49	3.2	3.13
D39599	37	44	0.49	7	6.93
D39020	581.7	590.2	0.46	8.5	7.42
D39177	347.7	355.2	0.46	7.5	7.47
D39593	115.6	143.6	0.45	28	26.25
D39548	16.5	18.5	0.45	2	1.95
D39357	756.5	761	0.45	4.5	3.38
D36105	217.3	218.8	0.45	1.5	1.43

- 105 - November 17, 2006

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D39126	490.5	500	0.44	9.5	9.40
D36156	45.5	94.3	0.43	48.8	3.90
D39552	296	315.7	0.43	19.7	19.33
D36021	190.2	193.4	0.42	3.2	2.66
D39291	101.7	106.6	0.42	4.9	4.82
D39031	533	540.5	0.40	7.5	4.82
D39175	542.3	546.3	0.40	4	3.54
D39009	772.8	781	0.39	8.2	6.17
D36014	1574.4	1576	0.39	1.6	0.88
D36039	211.3	215.6	0.39	4.3	3.04
D39020	601.6	610.7	0.38	9.1	7.96
D36044	193.3	195.9	0.37	2.6	2.24
D39012	515.1	517.2	0.37	2.1	1.70
D39180	595	598.1	0.37	3.1	3.07
D39595	77.9	88	0.36	10.1	9.93
D39158	303.5	324.5	0.36	21	17.96
D39155	180	186.2	0.35	6.2	5.11

Table 18.36   Project 308_27_56_561 - 561 Zone Significant Drill intercepts

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D30662	407.9	409.7	30.07	1.8	1.28
D27174	38.3	39.5	16.01	1.2	1.02
D28311	111.5	114.1	13.53	2.6	2.17
D28310	90.5	95.5	13.44	5	4.12
D30371	142	175.5	9.95	33.5	2.12
D27364	130.3	135.2	9.06	4.9	1.83
D30773	380.3	383.1	8.10	2.8	2.78
D30856	254.4	260	7.16	5.6	1.62
D30665	329.5	334.4	7.01	4.9	3.29
D28429	134.1	138.2	6.78	4.1	1.57
D27174	34.3	35.4	6.43	1.1	0.93
D27351	62.6	64.7	6.34	2.1	0.83

- 106 - November 17, 2006

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D30870	320.9	321.4	5.42	0.5	0.36
D27389	100.1	102.2	5.34	2.1	1.18
D27258	849.5	853.5	4.96	4	3.55
D271129	146.1	171	4.91	24.9	6.78
D28449	110	113.1	4.65	3.1	2.48
D28428	66.5	67.8	4.21	1.3	0.99
D27363	95	99.4	4.13	4.4	2.10
D30661	280.5	288.4	3.88	7.9	6.13
D27277	49.9	55.7	3.87	5.8	4.95
D27282	158.4	163.7	3.84	5.3	1.39
D30714	378.5	382	3.59	3.5	3.48
D30834	224.5	225.3	3.49	0.8	0.58
D28437	78.5	80.2	3.43	1.7	1.70
D30662	311.9	324.2	3.28	12.3	8.69
D30766	151	154	2.52	3	2.74
D30832	225	228.7	2.45	3.7	3.19
D271128	86.6	94.9	2.32	8.3	3.81
D27280	68.1	71.1	2.30	3	2.55
D271075	1.6	7.9	2.11	6.3	4.74
D30756	137.3	143.3	2.11	6	1.75
D27286	134.4	146.4	2.11	12	2.34
D27442	73.6	77.2	2.01	3.6	3.08
D28436	200	206	1.98	6	2.25
D27167	68.9	73.6	1.85	4.7	3.88
D28396	148	152.5	1.76	4.5	2.65
D271233	17.7	22.8	1.75	5.1	4.18
D28431	68.3	70.6	1.73	2.3	2.24
D28439	116.5	123.8	1.63	7.3	4.45
D271095	60.3	71.7	1.62	11.4	2.69
D30879	408.2	415.5	1.49	7.3	5.98
D30835	200.5	203.8	1.45	3.3	2.61
D30780	501.8	504	1.45	2.2	1.74
D27374	81.9	87.4	1.44	5.5	4.04
D30723	450	455.7	1.43	5.7	5.05
D30713	376	379.4	1.38	3.4	3.32

- 107 - November 17, 2006

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D30707	395.1	396.7	1.37	1.6	1.59
D27382	53.7	58.4	1.32	4.7	3.37
D27170	55.5	65.9	1.21	10.4	9.17
D28315	89	97.5	1.20	8.5	7.41
D271095	117.8	120.7	1.16	2.9	2.69
D28432	81.4	83.6	1.11	2.2	1.93
D30725	513.3	520	1.09	6.7	5.30
D28434	110.5	113.6	1.09	3.1	1.91
D27234	819.7	823.3	1.08	3.6	3.56
D27397	40.2	41.8	1.05	1.6	1.05
D30600	201	203.1	1.05	2.1	2.10
D27405	250.2	253.7	1.02	3.5	3.20
D27279	42.7	46.4	0.99	3.7	3.21
D28440	193	205.9	0.97	12.9	5.02
D30706	391	394.3	0.92	3.3	3.29
D28433	99.6	102.8	0.89	3.2	2.39
D28380	4	17.2	0.86	13.2	11.00
D27290	138.8	141.3	0.85	2.5	0.64
D27450	4.1	6.3	0.85	2.2	1.94
D271092	173.1	179.8	0.82	6.7	6.69
D27353	128.6	132	0.81	3.4	0.86
D28146	239.1	242.3	0.79	3.2	1.49
D30469	0	1.8	0.77	1.8	1.79
D27390	57.5	60.2	0.76	2.7	2.06
D27257	655.2	656.7	0.72	1.5	1.45
D271086	189.3	192.6	0.71	3.3	3.29
D2725	15.3	18.7	0.70	3.4	2.19
D30715	392.4	394.8	0.69	2.4	2.38

Table 18.37 project 42_36_DC - DC Zone Significant Drill Intercepts

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D39490	216.5	219	45.35	2.5	1.68
D39489	204	212	12.25	8	6.34

- 108 - November 17, 2006

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D40175	205.3	213	10.74	7.7	7.39
D39041	527.7	530.8	8.28	3.1	2.18
D39488	203	206.3	8.10	3.3	2.38
D39639	209	211.7	7.61	2.7	1.97
D39057	692.3	707.5	7.54	15.2	9.20
D39467	321.6	324.2	7.03	2.6	2.44
D40004	1977.3	1979.4	7.00	2.1	1.86
D39638	165.1	166.8	6.44	1.7	1.49
D39491	237.4	241	6.41	3.6	2.07
D40064	268	274.1	6.13	6.1	3.95
D40073	34.5	43.6	5.88	9.1	5.03
D39040	645.59	651.19	5.78	5.6	3.49
D40033	185.2	191	5.69	5.8	4.06
D40064	298.7	303.5	5.02	4.8	3.10
D40036	201	205	4.55	4	2.99
D39380	90	92.5	4.09	2.5	1.87
D40182	257	259.1	4.06	2.1	1.85
D39490	338.6	344.5	3.87	5.9	3.96
D39051	526	538.4	3.82	12.4	10.65
D39418	12.5	21	3.70	8.5	6.54
D39109	533.5	536.5	3.24	3	2.19
D39429	316	319.4	3.16	3.4	2.81
D39122	633.6	638	3.09	4.4	2.64
D40063	315.9	329.5	2.79	13.6	8.15
D39495	223	225	2.71	2	1.34
D40136	20.4	24.2	2.70	3.8	3.15
D40019	256.5	276	2.65	19.5	9.03
D40037	228	234.4	2.60	6.4	5.00
D38009	185	190	2.40	5	4.22
D39468	324.9	327.5	2.35	2.6	2.51
D39367	73	78.3	2.32	5.3	3.75
D39516	397.4	401.1	2.26	3.7	3.50
D40087	201.3	216.6	2.25	15.3	14.74
D39419	15	21.1	2.20	6.1	4.82
D40036	250	255.5	2.17	5.5	4.12

- 109 - November 17, 2006

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D40028	135	137.2	1.99	2.2	2.14
D39408	319	322.7	1.97	3.7	3.26
D39106	506.71	516.01	1.94	9.3	7.79
D39499	224.5	228.6	1.94	4.1	3.12
D40019	243.8	251.2	1.83	7.4	3.50
D40023	161.6	169.5	1.82	7.9	7.15
D40032	159	168.7	1.81	9.7	7.71
D39066	678.3	684.2	1.81	5.9	3.95
D39059	799.61	804.81	1.76	5.2	2.64
D39127	790.6	816	1.76	25.4	13.38
D39370	57.4	61.1	1.71	3.7	2.47
D39368	70.7	76.8	1.67	6.1	1.75
D40181	289.1	298.8	1.67	9.7	8.38
D39077	514.2	537.3	1.65	23.1	20.12
D39052	478.5	482	1.57	3.5	2.96
D39120	631	645.8	1.56	14.8	10.50
D39073	739.2	755.6	1.54	16.4	9.69
D39109	579.9	582.2	1.53	2.3	1.71
D40069	266.8	273.1	1.52	6.3	4.35
D40030	230	242.6	1.51	12.6	6.37
D39373	76.7	85	1.51	8.3	5.31
D40035	203.4	207.7	1.46	4.3	2.61
D40177	239	243.8	1.45	4.8	3.64
D39089	434.5	437	1.42	2.5	2.24
D40088	226.3	237.7	1.37	11.4	11.14
D39131	574.4	577	1.35	2.6	1.69
D40158	26	29.9	1.32	3.9	3.30
D40024	155.9	159.2	1.31	3.3	2.75
D39492	200	212	1.30	12	9.96
D39430	337.9	341.2	1.28	3.3	2.67
D40016	159	163.8	1.26	4.8	4.03
D39110	646.61	658.91	1.22	12.3	7.35
D39047	393.01	407.71	1.16	14.7	14.25
D40019	304	319.9	1.11	15.9	6.98
D39430	315	320.1	1.09	5.1	4.12

- 110 - November 17, 2006

Hole Number	From (feet)	To      (feet)	Gold Grade (oz/ton)	Length (feet)	True
					Thickness
					(feet)
D40066	368.7	383.7	1.07	15	7.68
D38009	180	182	1.07	2	1.69
D40086	239	252.1	1.04	13.1	11.70
D40177	266	284.5	1.04	18.5	14.12

19.0        OTHER RELEVANT DATA AND INFORMATION

19.1        Production Plan and Schedule

The details of the Strategic Business Plan (“SBP”) for the Red Lake Gold Mines are confidential. However, it is public knowledge that RLGM strives for the 1Moz/yr target. To reach this target RLGM will rely on the Campbell Complex, Red Lake Complex and satellite deposits in the district.

The hoisting and milling capacities are illustrated in the figures below. Plenty of hoisting capacity will be available with the #3 shaft connection to both underground complexes. A mill upgrade (1000 t/day) could be necessary to mine the sulphides & satellite deposits.

Figures 19.1 and 19.2 Hoisting and Milling Capacity for Combined Operations

- 111 - November 17, 2006

19.3.1     Cutoff Grade

Cutoff grade on site is defined as the break-even grade required to cover the cash costs. RLGM has used a mix of cut off grades & economic calculations to establishing the reserves.

Cut Off Grades Vs Complex

Complex	Average Cut off Grade (Oz/t)
Campbell Complex	0.35
Red Lake Complex HGZ	0.40
Red Lake FW sulphides	0.23

Each individual stope is evaluated economically. The geologists provide the block model to the engineers which create a mine design. The mine design is an assembly of development & stope wireframes and a sequence of mining with a schedule applied to it. The software allows the creation of different scenarios & mining methods to efficiently mine the resource dictated by the Block model. The Block model information provides the category of resources (Proven, Probable, inferred). With this information, every component of the design is costed and analyzed versus the revenue generated. If the stope is economical and passes the financial criteria, then the stope can be added to reserves.

19.1.1     Mining Extraction and Ore Losses

RLGM uses 100% mine recovery for scheduling the LOM plan reserves. Reconciliation from both mines indicates that such recovery levels can be met due to the high grade nature of the orebodies and the inferred ore around the stopes not calculated in the reserves.

- 112 - November 17, 2006

Figure 19.3 Reserves and Production for Combined Operations

19.2        Red Lake Complex

The Red Lake mine is serviced by two shafts, soon to be complimented by the #3 shaft, to be commissioned in January 2008. The new No. 3 Shaft was collared in late 2003, is underway to access new and deeper High Grade Zone Mineral Reserves and some of the lower grade Sulphide Zone Mineral Reserves. The No. 3 shaft will provide increased hoisting capacity, reduce access time to the work face, and increase ventilation facilitating more mining equipment and additional mining horizons.

The historic Red Lake Complex head frame is set over the mine’s current principal operating shaft (#1 Shaft) which extends to a depth of 1,023 metres and serves the mine from 1 Level to 23 Level. The first four levels are spaced 38.1 vertical metres apart, and the remaining levels are spaced 45.7 vertical metres apart. On 23L, a main access drift connects #1 Shaft to #2 Shaft (winze) which is situated 914.4 metres to the south and which services the mine from 23 Level to 38 Level. A branch of this main track access drift also connects to a main access ramp, which connects the levels from 21 Level to 42 Level. This ramp is planned to eventually reach 50 Level. The total current depth of the mine is over 1,829 metres. The shafts are connected to the ramp on the levels by 9 feet by 9 feet track drifts. The ore zones are connected to the ramp by 10 feet by 13 feet trackless drifts. The ramp is also 10 feet by 13 feet at 17% grade. The #1 Shaft has two production compartments, which are equipped with skip/cage and skip/counterweight conveyances operating in balance. The #1 Shaft also has one service compartment equipped with an eighteen person double deck cage for moving personnel and material. The #2 Shaft is equipped

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with two skip/cage conveyances operating in balance for handling production, personnel and material. The #2 Shaft cages can carry seven people.

Mining is carried out using predominantly underhand and overhand cut and fill techniques allowing maximum ore extraction and minimal dilution. The high-grade, narrow vein system is being mined at the rate of approximately 750 tons per day with an average grade of 1.5-2.0 ounces per ton (60 grams/ tonne). The high-grade mineralisation and complex geometry of the ore lenses require operating under tight geologic grade control..

Jacklegs, stopers, longtoms, two boom air jumbos, and single boom electric/hydraulic jumbos are used for drilling ore and waste. Mucking machines and trains, scooptrams ranging in size from 1 yard to 3.5 yard capacity, and 16 ton trucks are used to move the broken rock. The most common explosive used in blasting is ANFO. Ground support consists of various combinations of rebar bolts, swellex bolts, and screen, depending on the requirements of the heading being driven. Blasting is carried twice a day when all workers are out of the mine, and is initiated by an electrical central blasting system. On-shift blasting is heavily restricted, and only permitted with proper guarding procedures in place. Broken muck is dumped into passes located near the ore zones by trucks and scooptrams. The muck is pulled from chutes on 37 Level and 34 Level, and hauled by train to passes near #2 Shaft, which feed the loading pocket on 38 Level. At the loading pocket, muck is loaded into the 6 ton skips, and hoisted up #2 Shaft, and dumped into bins on 23 Level. A 36 ton train is used to transfer the muck across 23 Level to passes that feed another loading pocket at the bottom of #1 Shaft. From that loading pocket, muck is loaded onto 4 ton skips and hoisted up #1 Shaft, and dumped into bins on surface that either feed the process plant or the trucks that haul to the waste dump.

The ventilation system is a push-pull design, with intake and exhaust fans on surface, and booster fans underground on 22 Level and 30 Level, producing airflow of 150,000 cubic feet per minute of fresh air reaching the High Grade Zone. Many drifts, raises, and ramps, plus the two shafts, make up the main ventilation circuit. Auxiliary fans of varying sizes bring the fresh air from the main ventilation circuit to the working faces.

19.2.1     Ground Conditions

The Red Lake mine is situated on the hanging-wall side of a southeast-plunging anticline, within the eastern portion of the Red Lake greenstone belt.

The dominant rock type in the mine consists of tholeiitic to komatiitic flows. These rocks are pillowed, often amygdaloidal, or massive, and where unaltered, mafic flows are dark green to black and aphanitic to fine-grained. Mafic volcanics, the main host for ore at the mine, are

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banded in appearance and moderately to strongly foliated. Diorite, which occurs as thick flows or sills, is a coarse-grained equivalent of these basalts.

Intercalated with the mafic volcanic rocks is a highly carbonatized and altered unit, which is believed to be an altered ultramafic rock of either volcanic or plutonic origin. This unit varies considerably from rhyolite breccia to talc-chlorite schist, to carbonatized andalusite-rich metasomatized rock. This unit is characterized by intense shearing and alteration, as well as being relatively incompetent.

Interbedded interflow sulphide-facies banded iron formation is intercalated with the mafic volcanic rocks. In most locations this unit is folded about a northwest-trending axial plane with the main bedding attitude striking east-west, and dipping vertically.

Above this package of mafic rocks are felsic flows, pyroclastic, clastic and chemical sedimentary rocks. Rhyolitic flows are mainly seen west of the main workings of the mine, while felsic volcanic breccias found near No. 2 shaft on 35 level appear to grade into lapilli-tuff on surface. Greywacke and chert occur interbedded.

Rockmass Strength

Strength and deformational testing have been carried out over the years at the Red Lake Mine. The Red Lake Mine Basalt and Altered Ultramafic rock types have a wide range of uniaxial compressive strengths. For modelling purposes, a UCS of 180MPa is used.

Rockmass Structure

The volcanic and sedimentary rocks and ore zones have been intruded by quartz-feldspar-porphyries (QFP), metadiabase, peridotite/serpentinite and lamprophyres.

Faults and shear zones were developed contemporaneous with folding, resulting in block movement along a mainly northwest-southeast trend. Much of the movement was of a cyclical nature, constantly brecciaing and resealing the structure before and during ore body formation.

There has been one (1) reportable fall of ground since mining began in the HGZ. It was related to unfavourable joints. The joints sets found in the HGZ are summarized in Table 19.1.

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Table 19.1:    HGZ Joint Sets

Joint Set	Dip	Dip Dir.
JS 1 (fabric)	70	225
JS 2	30	095
JS 3	40	015
JS 4	65	260

Stress

Subsequent to review of the general ore trends below 30Level and employing principal stress directions perpendicular and parallel to the ore body the following have been employed for purposes of induced stress determination: (Smith, 1987, Arjang/Harget, 1991)

Figure 19.4: HGZ Principle Stresses

HGZ Reportable Rockbursts

There have been three (3) reportable rockbursts (> 5.5 tons displaced) since mining began in the HGZ. The ground support system has contained the displaced material from all other events.

Typically, macroseismic events damage the rock causing it to bag in the screen, while the bolts remain intact, holding the screen. This loose must then be removed from the screen such that the system is no longer pre-loaded (in case of a second seismic event). The paste appears to behave much better under seismic loading than the rock:  There has yet to be damage in an UCF stope due to seismic activity, despite significant damage to nearby rock.

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19.3Access Development

The Red Lake mine is accessed from the footwall side of the ore body by #2 Shaft. The mine operates two hoist shafts. The #1 shaft extends to just below the 23^rd level at approximately 3590 feet below surface, while the #2 shaft (an internal winze) begins at the 23^rd level and extends to below the 38^th level at 5700 feet below surface. The main levels have been driven from the shafts at approximately 150-foot intervals.

The existing development for production mining is from 30 level down to 41 level. Stope accesses are driven from each sub-level to the different ore lenses. End and central stope accesses are used depending on the location of the ore lens relative to other lenses and sequencing requirements. Stope access dimensions are determined by the target ore width and ventilation requirement. Stope access gradients typically range from –20% to +20%. Ore width determines the size of the load haul dump (LHD) units used for mucking stopes. The LHD’s range in size from 1yd³ to 3.5 yd³.

The Red Lake Complex is in the process of completing a new 1,924 metres (6,312 foot) deep shaft to access ore reserves at depth. Sinking of this 21-foot diameter shaft is expected to complete 4th quarter 2006. It is anticipated that hoisting could begin in the 3rd quarter 2007, and overall completion of the project in December 2007. The new shaft will allow for increased hoisting capacity, reduced travel time to the High Grade Zone, and act as a path for fresh air once the ventilation upgrade is complete.

19.1.3Mining Method

The mining methods currently in use at the Red Lake Complex can be grouped into 4 broad categories. They are: Overhand Cut & Fill (OCF), Underhand Cut & Fill (UCF), Pillar Recovery, and Longhole. Table 19.2 below summarizes the ratio of break tons by mining method.

Table 19.2: RLC Mining Methods

	2003	2004	2005	2006E
Overhand Cut & Fill (OCF)	66%	54%	52%	37%
Underhand Cut & Fill (UCF)	29%	42%	43%	40%
Pillar Recovery	5%	4%	5%	8%
Longhole	-	-	-	15%

The methods were chosen for the following reasons:

1.To selectively mine highly variable and complex ore structures, allowing for full production geology control of each ore blast. In most cases, sufficient unknown ore is found during mining to cover the mining costs.

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2.To reduce dilution in order to make the most of the limited available hoisting capacity by optimizing the grade of the ore hoisted.

3.To mitigate the potential for and damage from seismic events by controlling the open mining span, providing better control of the mining sequence, and minimizing the creation of sill pillars.

4.To provide the number of mining fronts required to continue to produce 700 plus tpd of ore, as the initial overhand cut and fill mining fronts reached sill position, without creating additional pillars.

5.To maximize ore recovery.

6.To develop methods that could be applied to future mine expansion at depth.

Overhand Cut & Fill (OCF)

Until recently, the majority of HGZ stopes were mined using OCF. Our current 2006 mining plan estimates approximately 37% of production to come from OCF stopes.

In OCF mining a horizontal slice of ore is excavated (~12ft high at RLC) and filled. The next cut is then mined above the filled cut. Whenever possible waste rock is placed as fill and capped with pastefill.

Most often, OCF stopes are accessed via a ramp. The ramp provides access for mechanized equipment. This method is sometimes referred to as Mechanized OCF. Captive OCF mining is limited to ore lenses that are isolated or small.

In areas where ore widths exceed maximum allowable spans, panel mining is instituted. Panel mining is a modified cut and fill technique where a fraction of a wide cut (a panel) is mined and filled tightly or braced with shotcrete pillars to support the back. Additional panels are then mined alongside the filled panel or panels as necessary, until all the ore is removed from the cut. The panel widths are chosen based on ground conditions.

Underhand Cut & Fill (UCF)

In early 2003, as some mining fronts moved into pillar recovery mining, the underhand cut and fill method was introduced. After a cut is mined, an engineered fill mat is placed in the stope before filling with pastefill. The next cut is then taken under the fill mat. All current UCF stopes are accessed via ramps. In UCF, benching of the ramp provides access to the cut below.

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Some of the benefits of UCF mining include greater wall stability resulting in lower dilution and less potential for damage from seismic events. UCF mining is the lowest cost mining method by a slight margin.

Paneling (as described in OCF mining) is also practiced in UCF mining when ore width exceeds maximum UCF mining widths based on fill mat design.

Pillar Recovery

Horizontal (sill) pillars are created between advancing mining fronts. These pillars, usually 40-60 ft thick, can pose significant mining and safety risks due to higher stress conditions resulting in lost productivity and added ground support costs. Thus, the mining method for a pillar is often selected and engineered to suit the unique conditions of that pillar. Several mining methods are utilized currently to recover ore from horizontal pillars.

Stope Sequencing

Stope sequencing is adjusted where possible in an attempt to reduce the risk of macroseismic activity induced by stress.

Due to the apparent benefits under seismic conditions, it is planned to use UCF as the primary mining method below 37 level. OCF will only be used to mine the first four to five cuts on a new mining horizon to set the required pastefill plug (minimum 40 vertical feet) before starting UCF mining.

19.1.4Equipment

The current mobile diesel equipment fleet consists of scoop trams, dump trucks, mine service and personnel vehicles, jumbo drills, bolting platforms, scissor lifts, tractors, forklifts, boom trucks, and utility trucks. A detailed equipment list is provided in Table 19.3.

As the expansion project comes to completion and mining progresses deeper, the equipment fleet will change accordingly. Capital has been budgeted for equipment additions, replacements, and rebuilds.

In addition to the mobile fleet, other equipment used includes: air and electric slushers, primary ventilation fans, dewatering pumps, rockbreakers, shotcrete machines, rail cars, rail mancars, forklifts, loaders, cranes, and tuggers.

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Table 19.3 Mobile Equipment List for Red Lake Complex

Type of Equipment	Make/Model	Quantity	Ownership
Production Equipment
3yd - 3.5yd Scoops	EJC 115 Scooptram	5	2 Goldcorp, 3 Contractor
	EJC 130 3.5 Scooptram	2	Contractor
	EJC 145 E Scooptram	2	Goldcorp
1.5yd - 2.5yd Scoops	JCI 125 1.5 Scooptram	5	3 Goldcorp, 2 Contractor
	JCI 125 E 1.5 Scooptram	1	Goldcorp
	EJC 100 E 2.5 Scooptram	1	Contractor
	Wagner 2G Scooptram	3	Goldcorp
Jumbos	MJM20B Pneumatic Jumbo	3	Contractor
	HS 107C Monomatic Jumbo	1	Contractor
	HS 105-40 Monomatic Jumbo	1	Goldcorp
Longhole Drills	Boart BCI-2	2	Contractor
	Boart Bar & Arm	2	Contractor
Dump Trucks	EJC 416D Dump Truck	2	Contractor
	EJC 415D Dump Truck	1	Contractor
	EJC 417D Dump Truck	2	Goldcorp
Bolting Platforms	Timberjack Bolting Platform	3	Contractor
	Teledyne Scissor Lift	2	Contractor
Muck Machines	Atlas Copco LM56	10	5 Goldcorp, 5 Contractor
	Eimco 12B	1	Goldcorp
	Eimco 21B	1	Goldcorp
Long Toms	2 Boom	2	Goldcorp
	3 Boom	2	Goldcorp
	3 Boom Rubber-tire	1	Goldcorp
Alimaks	STH5 Alimak	3	Contractor
	Alicab	1	Contractor
Ancillary Equipment
Track Motors	Balco 4 Ton Battery	2	Goldcorp
	Mancha Battery	6	3 Goldcorp, 3 Contractor
	Atlas 3 Ton Battery	8	Goldcorp
	Clayton 4.5 Ton Battery	5	Contractor
	Warren Nex-Tec Battery	5	Goldcorp
	Clayton 6 Ton Battery	1	Goldcorp
	Goodman 8 Ton Diesel	2	Goldcorp
Utility Vehicles	Elmac D5 - 4A	3	Contractor
Boom Trucks	JCC 8 Boom Truck	1	Contractor
	Teledyne LP15ARN	1	Contractor
Personnel Vehicles	Miller Minekart	1	Goldcorp
	Miller Mine Kart 4x4	7	Goldcorp
	Minecat 108 PC	1	Contractor
	Kubota RTV 900	5	Contractor
Tractor	Kubota L2410 DT	1	Contractor
Forklift	Manatou BT420	1	Contractor
Backhoe	Kubota R520 Minemaster	1	Contractor
	Kubota R410 Backhoe	1	Goldcorp
	Teledyne Pneumatic Backhoe	1	Goldcorp

19.4Infrastructure

19.4.1Site Facilities

The Red Lake Complex’s process plant, pastefill plant, administration buildings, maintenance shop, warehouse, coreshack and camp are located at the mine site. The camp can accommodate upto 176 personnel at one time and is currently at capacity. It is equipped with accomodations,

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cooking, and dining facilities. Personnel who stay in camp have access to Campbell Recreational Centre (at the Campbell Complex). The camp houses most contract employees. The remaining employees and staff live in town.

19.4.2Power

Electric power to the Red Lake Complex is supplied from a 44 kilovolt transmission line from the Red Lake Transformer Station. The main incoming line splits at a metering tower, with two legs supplying a substation for surface and underground, and another high voltage line feeding a substation for the process plant.

Current demand for the Red Lake mine site is approximately 12 megawatts; however, during the winter heating season, demand has peaked at slightly over 13 megawatts. The process plant consumes about 4 megawatts of power, with the remainder being consumed by the underground and the surface plant buildings.

19.4.3Ventilation

The present Red Lake Complex ventilation network utilizes a “push / pull” system through a complex array of interconnected raises, abandoned stopes, boreholes, chutes, ore/waste passes, drifts and ramps. Because of the history of development, the mine does not have a direct and dedicated intake and exhaust route to and from the current production areas in the mine. The network integrates a multitude of parallel routes between most levels to reduce resistance for ventilating the mine. In combination with the main surface fresh air fan and the surface exhaust fan, underground booster fans are used to direct the airflow to lower areas of the mine. Some additional fresh air is being drawn off the #3 Shaft on 37-1Sub down a dedicated raise to 42 Level.

The mine fresh air is provided by a surface downcast fan generating approximately 165 kcfm. This air travels through a variety of paths, as described above, until it reaches 30 Level. At this point the fresh air is passed through the 30 Level Booster Fans providing 150 kcfm to the main production areas.

The mine currently maximizes the available airflow for the operation of the diesel fleet. A major ventilation upgrade is in progress which will significantly increase the capacity of the system.

19.4.4Dewatering

The main mine dewatering system comprises 5 main sumps on 37, 30, 23, 15, and 8 levels. The sumps range in size from 105,000 gallon reservoirs on 23 level to 70,000 gallons on 30 level.

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Each main sump has a clear and ‘dirty’ water side. The various sub-levels report through mine drainage systems via air powered diaphragm pumps, drainage boreholes, or small submersible pumps in the case of the ramp.

Water demand is approximately 201,000 U.S. gallons per day. Ground water infiltration amounts to about 75,000 U.S. gallons per day primarily through the upper levels. The result is approximately 275,000 U.S. gallons per day is pumped from the mine. The discharge water reports to the paste plant to satisfy backfill requirements. Any unused water is discharged to the tailings impoundment area.

19.4.5Compressed Air

The mine air compressor plant comprises 6 Sullair, two stage tandem screw type compressors. Each is powered by a 450 horsepower, 600 volt motor. Each compressor is rated at 2200 cfm at 135 psi. The compressors are housed on surface in a separate stand alone building. The compressors discharge into a common 8” header, and then into the main distribution pipeline located in a borehole, shafts, then through the mine network. A small amount (approximately 1000 cfm) is sent to the process plant and a lesser amount is sent to the paste plant.

Demand is controlled by a ‘Supervisor 2’ logic system, and as demand increases another compressor comes on line. There is always one compressor in ‘trim’ mode to allow for quick reaction to demand changes.

19.4.6Communication

Communication is enabled throughout the mine via leaky feeder and/or Bell Canada telephone system. There is also a fiber optic network underground used for ventilation and seismic monitoring, controls, and computer network access to the underground office.

19.5Campbell Complex

Shaft sinking began in 1946 with production commencing in 1949 at a rate of 300 tons/d, from seven levels at 150 ft (45 m) spacing. Since then the Campbell shaft (or No. 1 Shaft) has been deepened four times and the four-compartment shaft was completed to below 27 Level, a depth of 1,316 metres below surface. There are 27 Levels at 45 metre vertical intervals, with an average of 6,000 metres of development per level. In 1999, the Reid Shaft was built to open up access to the deep underground zones including the DC Zone.. The Reid Shaft is located 150 metres west of the Campbell Shaft, and extends to a depth of 1,819 metres. In 2003 development began on the DC, or Deep Campbell, zone. Capital development work continued in 2006 with spiralling ramp development from the bottom of the shaft at 39 Level down in order to open up the DC Zone

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between 41 and 43 levels. In 2006 the 4199 Ramp was driven into the hangingwall to the DC zone and a 400 m ‘T’ shaped drill platform completed. This drill drift is positioned to test the Deep Campbell depth extent an additional 600 m below 41 Level.

Above 27 Level, a combination of mechanized, rubber-tired diesel equipment and conventional track haulage is utilized for mining. Full track haulage facilities exist on all 27 levels. Below 27 Level, all mining is mechanized to provide greater flexibility and productivity. In 2005, access between 36 level through to 39 Level was achieved via an incline from 39 level. The decline down from 27 Level is scheduled to reach the 33 Level shaft station by 2006 year-end.

In 2005 a total of 718,000 tonnes were mined, of which 440,000 tonnes (61%) was ore. Of the ore mined, 73% was obtained from long hole stopes, 17% from cut and fill stopes and 10% from ore development. In 2005 the quantity of ore sourced from cut and fill stoping increased and this will continue in 2006. Cut and fill stoping will help improve the quality of ore in areas where dilution is high and the orebody is narrow or discontinuous. The mine will continue to focus on reducing seismic risk and dilution in long hole stopes by minimizing stope spans, decreasing the size of the blasts and continuing to fill all stope excavations with high quality paste fill. In 2005, 35% of mill tailings were sent underground as paste fill.

19.5.1Mine Plan

Mine access is through two separate shafts. The No. 1 (Campbell) Shaft is a four-compartment shaft sunk to below 27 Level, a depth of 4318 feet below the surface. There are 27 Levels at 150 foot vertical intervals, with an average of 19,700 feet of development per level.

The Reid Shaft is located 500 feet west of the No. 1 (Campbell) Shaft, and extends to a depth of 5,968 feet with access to 7, 17, 27, 30, 33, 36, 39 levels and the 40 loading pocket.

Above 27 Level, a combination of mechanized, rubber-tired diesel equipment and conventional track haulage is utilized for mining. Full track haulage facilities exist on all 27 levels. Below 27 Level, all mining is trackless to provide greater flexibility and productivity. Ramp access (-15%) is provided from 27 to 33 levels and from 36 to 41 levels. Capital development work continues to open up the DC Zone between 36 and 43 levels.

Ore is crushed underground, hoisted to the surface via the Reid Shaft, and then re-crushed to below 10mm size. Completed stopes are backfilled with cemented pastefill.

The gold production target for this operation is 240,000 oz to 255,000 oz/a over the life of the mine.

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19.5.2Geology Summary and Ground Conditions

Campbell is located in the eastern part of the Red Lake Greenstone Belt, in the Birch Lake/Uchi Lake sub-province of the Canadian Shield. Gold veins average 1.5 - 1.8 m in width and extend over strike length of 30 to 300m. m.

Sulphide replacement zones vary from 3-12 m in thickness and extend over a strike length of 120 to 180 m. Gold occurs as either free gold or is encapsulated with sulphide minerals, mainly arsenopyrite, pyrite and pyrrhotite. The arsenopyrite is a marker for gold ore. A minor amount of silver occurs with the gold.

There are six basic rock types at Campbell mine, excluding the ore zones, these include andesite, ultramafic, rhyolite, Lamprophyre dyke, altered dyke, Campbell diorite.

Rockmass Strength

The material properties of each rock unit excluding the altered dyke and the Campbell diorite are listed in the following table. For the altered dyke, the ultramafic rock properties are utilized.

Table 19.4 Material Properties of Rock Units

Rock Type	UCS	Tensile Strength	Young’s Modulus	Poisson’s Ratio
	(MPa)	(MPa)	(GPa)	(v)
Andesite	180	16	69.9	0.21
Rhyolite	197	-	89.3	0.19
Ultramafic	126	4.8	56.7	0.19
Lamprophyre	188	16	43.4	0.20

Rockmass Structure

Since 1997, Campbell Mine’s Geology Department has being working on defining the mine-scale, major structures and lithologies. This information has been compiled into a digital model based on 100-scale digitized geology plans. The major faults are described below.

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Table 19.5 Major Faults at Campbell Mine

Fault	Orientation	Description
19	270-280 / 70-85SW	Breccia & gouge; not dyke
SXC	315 / 50SW	Wide (10-30 ft) breccia, gouge & black line fault zones; local QFP dyke
West Drift	300-345 / 45-58SW	Wide (up to 100’) breccia, gouge and black line fault zones; sheared 5A dyke common; rhyolite slivers common
Campbell &	300-315 / 65-90SW	Ductile shear zone >10’ wide with black line faults
East Campbell
70	300-315 / 70SW	Tight black line fault zone on rhyolite-basalt contacts; thin breccia and gouge zone on komatiite-basalt contacts
51	310 / 55SW	Wide (10-30 ft) of breccia, gouge and black line fault zones; 5A or Lamp (5L) dyke on or close to fault plane
25 System	300 / 90	Thin breccia, gouge and black line fault zones; not dyke; sub-horizontal slickenside common
02	315 / 75-80SW	Thin breccia, gouge and black line fault zones; not dyke
04	295 / sub-vertical	Thin breccia, gouge and black line fault zones; not dyke
09	315 / 65SW	Locally thick quartz-filled breccia lenses mark fault

Mine Wide Joint Characteristics

The table below summarizes the mine-wide joint sets based on geotechnical mapping. In general, the joints are persistent, smooth to slightly rough, planar and tightly closed. The joint infillings vary from no infilling to surface staining to calcite, quartz, talc and sulphide infillings. The average residual friction angle for joints in the andesite, ultramafic and rhyolite units is 29°  (Bawden & Tod, 1996).

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Table 19.6 Joint Sets Identified at Campbell Mine

Joint Set	Strike / Dip	Description
JSA1	100-125 / 70-82	•Foliation
		•Typical joint spacing: 1 – 2 ft
JSA2	130-150 / 62-72	•Foliation
		•Typical joint spacing: 1 – 2 ft
JSB1	225-245 / 15-40	•Flat-lying joint
		•Typical joint spacing: 2 – 4 ft
JSB2	340-025 / 10-30	•Flat-lying joint
		•Typical joint spacing: 2 – 4 ft
JSC	020-050 / 60-85	•Cross-cuts foliation (common)
		•Typical joint spacing: 6 – 12 ft
JSD	350-010 / 55-75	•Random set (less common)
JSF	300-315 / 60-75	•Random set (less common)

Stress

A series of in situ stress measurements have been made at Campbell mine (See below). All of the measurements have been made using CSIR triaxial strain cells and all measurements have been made in the Andesite unit.

Table 19.7 In-situ Stress measurements for Campbell Mine

Depth	1	Orientation	2		3		E		Rock	Source
(ft)	(Mpa)		(Mpa)		(Mpa)		(GPa)		Type
1903	26.2	078/05	16.9	344/38	2.6	174/52	74.3	0.25	And	CANMET
1903	30.8	074/08	18.6	340/26	5.1	179/62	74.3	0.25	And	CANMET
2051	23.2	195/13	9.5	086/55	7.5	293/31	91.7	0.21	And	CANMET
2051	22.2	074/35	15	183/15	9.2	300/45	91.7	0.21	And	CANMET
2051	23.8	067/02	15.7	336/18	10.2	162/72	91.7	0.21	And	CANMET
2051	40.8	217/15	17.9	042/75	9.5	307/01	90.5	0.23	And	CANMET
2051	24	141/15	16.1	046/17	9.8	270/67	90.5	0.23	And	CANMET
2051	24.5	056/38	17.2	258/49	3.9	155/11	86.5	0.21	And	CANMET
2198	62.2	317/24	30.6	058/23	16.6	185/55	83.5	0.2	And	CANMET
2198	49.9	278/14	21.3	042/65	2.6	183/20	83.5	0.2	And	CANMET
2198	41.7	091/09	25	190/44	10	352/44	83.5	0.2	And	CANMET
3248	49.8	058/17	26.2	321/23	10.9	180/61	85.3	0.22	And	CANMET
3248	41.6	086/10	19	351/26	15.3	185/62	85.3	0.22	And	CANMET
3248	68.6	079/20	25.5	333/37	7.2	185/47	85.3	0.22	And	CANMET
4003	81.6	254/01	40.7	345/31	14.1	162/59	84.8	0.2	And	CANMET
4003	67.4	246/27	39.3	341/11	30.1	092/61	84.8	0.2	And	CANMET

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Depth	1	Orientation	2		3		E		Rock	Source
(ft)	(Mpa)		(Mpa)		(Mpa)		(GPa)		Type
4003	74.1	238/30	43.5	336/12	32.6	085/57	84.8	0.2	And	CANMET
4003	68.9	085/05	41.6	178/28	24.4	347/61			And	AECL
4003	68.2	239/25	43.8	330/02	21.1	063/65			And	AECL
4003	42.8	063/12	30.9	156/14	16.7	293/72			And	AECL
4003	69	297/09	49	206/06	37.1	085/79			And	AECL
4003	56.8	118/07	35.5	009/27	21.5	245/79			And	AECL
5400	81.6	235/34	60	129/22	23.9	013/47			And	CANMET
5850	80.5	235/11	68.6	132/47	38.4	335/41			And	CANMET

Based on these measurements, the stress tensors shown below have been adopted for mine-wide numerical modeling. However, field observations indicate that the in situ stress field rotates and is locally disturbed near lithologic contacts (e.g. dykes, Andesite-Ultramafic contacts, etc.).

Structures, such as the Andesite-Ultramafic contacts, are believed to play an important role in stress distribution. In general, the Andesites are more competent and bursting tends to concentrate around these contacts, with the seismicity clustering preferentially in the Andesite, as opposed to the Ultramafic units.

Table 19.8 Mine Wide Stress Tensor Campbell Complex

Principal Stress	Magnitude (MPa)A	Orientation
1	8.18 + 0.0129 * D	070/00
2	3.64 + 0.00841 * D	160/00
3	0.007 * D	000/90 (vertical)

^A Where D is the depth below surface in feet.

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19.5 Campbell Complex Stress Measurements as a Function of Depth

Reportable Ground Incidents

Incidents of ground instability (rockbursts exceeding 5 t and falls of ground exceeding 50 t) are reported in accordance to the Occupational Health and Safety Act. Historical frequencies of such incidents are outlined in the following graph.

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Figure 19.6 Reportable Ground Incidents

19.5.3Access Development

Access to the underground operations are achieved through the No. 1 (Campbell) shaft with access to all levels from 1-27 and the Reid Shaft with access to levels 7,17,27,30,33,36,39 and 40 loading pocket. Below 39 level, a decline is the primary access for all personnel, equipment and materials. Emergency egress is achieved through a ladder way from 41 to 39 level. From 39 level to 30 level, egress is achieved through an 8’ diameter borehole with an emergency egress hoist/cage. Upon completion of the ladder way between Campbell complex 36 level to Red Lake complex 34 level, this egress hoist will be decommissioned. The upper portion of the mine (above 33 level) achieves secondary egress through the ladder way in the No. 1 (Campbell) Shaft.

19.5.4Mining Method

The primary mining method at the Campbell Complex is longhole (65%) followed by conventional overhand cut and fill (30%) and approximately 5% development ore. The longhole mining includes both longhole panel mining and longhole sill recovery. Sill recovery in mined out (shrinkage) areas is done from footwall access drifts to minimize worker exposure to highly stressed areas within the ore zone. Longhole panel mining is performed on 50 foot sublevel intervals with typical widths of 8 to 20 feet and typical strike lengths from 50 to150 feet. Larger

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stopes may be mined transversely through the use of drawpoints on 50 foot intervals. Mechanised cut and fill is used over conventional longhole where feasible, especially in areas where the dilution is high and orebodies are narrow or discontinuous. Cut and fill headings are mined approximately nine feet high at widths ranging from 4 to 30+ feet depending on the ore zone. When longhole is the chosen method, stope spans are minimized, blast size is kept small, and pastefill is placed as soon as possible after each stope is mined out.

19.5.5Equipment

The current mining fleet is a combination of conventional and mechanized mining equipment. The fleet consists of many locomotives, muck machines, Cavos, Jacklegs, Stopers, Longtoms, single and twin boom Jumbos, scissor lifts, maintenance/service vehicles, grader, forklift, 1 to 4 yd³ scoop trams, and 16t haul trucks.

Board Longyear performs the longhole drilling at the Campbell Complex utilizing their own longhole drilling machines.

The table below shows a list of the major mobile equipment in the mine.

Table 19.9 Mobile Equipment at Campbell Mine Complex

Type	Model	Quantity
Locomotive	Atlas Loco	11
	Mancha Loco	19
	Balco Loco	15
	Warren Loco	14
Scoop	Toro 151D	5
	EJC 61D	2
	EJC 60E	9
	EJC 90D	1
	EJC 60D	1
	EJC 61E	1
	Toro 151D	4
	Toro 151E	5
	Wagner ST3.5	1
	EJC 100D	1
	EJC 65D	1
	Toro 150E	1
	Toro 006	5

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Type	Model	Quantity
Haul Truck	EJC 417	4
	Wagner 416	3
	Wagner 417	1
Jumbo Drills	Quasar NVD4E60	3
	Axera DO6 240	1
	Tamrock H105D	1
	Axera D06 226	1
Specialty/Service	Getman A64 Scissorlift	1
	Marcotte M40 Scissorlift	3
	Marcotte Shotcrete	1
	Kubota Forklift	1
	Miller 6500 Forklift	1
	Caterpilar 120G Grader	1
	Wagner 408 Mancarrier	1
	Tamrock Bolter	1
	Toyota Mech/elec	3
	Marcotte Shotcrete	1
	Miller 6500 Forklift	1
	Marcotte Grader	1
	Wagner 408 Mancarrier	1
	Kubota Mine Master	4
	Kubota Mine Master Mech	1

Fixed equipment and facilities are also typical for the planned mine layout and include primary ventilation fans, mine air heaters, dewatering pumps, explosive magazines, maintenance shops, fuelling stations and personnel refuge stations/lunchroom.

19.5.6Mining Dilution

Planned dilution has been accounted for in the mining plan. Planned dilution is incorporated into the mine plan and mucking schedule. Reconciliation is performed after each stope is mined in order to determine the mining effectiveness. Unplanned dilution may occur due to a number of factors including blast hole deviation, over loading, etc. The goal of the reconciliation process is to identify areas of significant dilution and rationalize why the dilution happened. With this information, the mine plan is modified where necessary based on lessons learned to improve the future product.

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19.5.7Infrastructure

Site Facilities

The Campbell Complex includes administration buildings, maintenance shop, warehouse Mill, Campbell and Reid Shaft hoist rooms, and core logging facilities. Potable water is supplied by the municipality. Process water is taken from Balmer Lake and Sandy Bay.

Power

Ontario Hydro supplies power to the Campbell Complex through lines from Vermillion Bay to Ear Falls and on to the Red Lake area. The Campbell Complex utilizes 14MW of power currently and the Red Lake Complex uses an additional 14MW. Provisions have been made to provide as much as 46MW total power, meaning excess capacity for future requirements. If further power is required in the future, a transformer upgrade at the Red Lake station would be required. There are multiple transformers throughout the site, which step voltage down to 4160, 2300 and 575 volts.

Diesel-powered generators provide the primary source of emergency power in the event of a disruption. The generators are intended as temporary power source to allow the mine site to maintain a reduced production rate until Ontario Hydro power is restored.

19.5.8Ventilation

The ventilation supply volume to Campbell mine is 500,000 cfm. Air is drawn into the mine primarily by exhaust fans. The system has 2 intake fans on surface with 1 intake underground booster fan and 4 surface exhaust fans with 7 exhaust underground booster fans. The surface intake fans are both equipped with propane heaters for winter operation however only 1 heater is required to operate at their current duties. The total fan BHP is 3200 HP (675 HP on intake) and all fans are equipped with VFD’s with normal duty at 1/2 to 2/3 of total BHP. The Campbell ventilation system has 3 circuits and each are primarily independent of one another. The upper circuit supplies 250,000 cfm between levels 27 and surface, the middle circuit supplies 150,000 cfm to levels 27 to 33 and the bottom circuit supplies 100,000 cfm to below 33 level. Ongoing work in 2006 & 2007 is designed to increase the total mine supply to 550,000 cfm with a 150,000 cfm increase to the bottom circuits and reduction on the upper circuit to accommodate for increased mining to depth.

19.5.9Dewatering

Mine water is handled by five underground pumping stations located on 7, 14, 21, 27 and 40. These stations pump approximately 53,000 ft3 of water daily. All underground water is recycled as mill process water.

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19.5.10Compressed Air

Three surface screw compressors deliver the necessary pressure for distribution underground through a network of pipelines connected through both shafts. Compressed air is used when drilling, and to power small hand-tools.

19.5.11Communication

Communication is enabled throughout most active mine headings by leeky feeder radio. Telephone service is provided in all shaft and refuge stations. Underground fiber optics networks are also utilized to transfer data from remote sensors and video feed from required areas.

19.6Environmental and Socio-economic

Red Lake Gold Mines operates under Goldcorp’s sustainability policy, which commits the operation to a defined standard of environmental stewardship. Sustainability is an important issue for every department. This involves protecting human health, reducing the impact of mining on the ecosystem(s), and returning the site to a state compatible with a healthy environment. Red Lake Gold Mines has developed a series of management programs for environmental activities, tailings management and occupational health and safety that enable the company to reach its commitments.

19.6.1Permitting Status

All requisite permits have been obtained for the mining and continued development of the mine site and are in good standing. The permitting application process has been very active in the last several years with the various expansion projects and process enhancements that have been underway at the 4 mine complexes. Attached below is a reference list of environmental permits for the mine.

The list below (Table 19.10) summarizes the active list of permits, licenses, certificates and registrations primarily for the Ministry of the Environment. The only documents not listed here are the Closure plan lists which are addressed in section 19.5. This list is divided into sections for the four active mine complexes which comprise the RLGM.

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Table 19.10: Environmental Permit and Approval List for Red Lake Gold Mine

GOLDCORP CANADA LTD. Red Lake Gold Mines
Permit List
Original Date: November 2, 2006            Update:
CAMPBELL COMPLEX		AGENCY	ISSUE DATE	EXPIRY  DATE
1.	COA – Industrial Sewage # 4-0058-86-928 Revisions No.1-4, Notice 1, 5, 6	MOE	July 1992	N/A
2.	PTTW # 6861-6R2JQ7 South Dam Lift Station	MOE	June 23, 2006	December 1, 2015
3.	PTTW # 7205-6R2GWS Balmer Lake Pump house	MOE	June 23, 2006	October 5, 2015
4.	PTTW # 5533-6R2GSM McManus Bay (Sandy Bay)	MOE	June 23, 2006	March 31, 2009
5.	Certificate of Approval (Air) #5342-5HKQ5U Comprehensive Air Permit	MOE	March, 2003	N/A
6.	Certificate of Approval (Air) #8-6020-95-006 Induction furnace, Baghouse	MOE	July 31, 1995	N/A
7.	Certificate of Approval (Air) #8-6001-91-006 Vent filter, Scrubber, Boiler	MOE	May 24, 1991	N/A
8.	Certificate of Approval (Air) #9416-666RTX Baghouse, Scrubber, Heaters	MOE	January, 2005	N/A
9	Certificate of Approval (Waste) # A860841 generator # ON0434400	MOE	May 11, 1994	N/A
10	Aggregate Permit #8816	MNR	July 24, 2006	N/A
11	Nuclear Permit # 11896-2-11-1	CNSC	September 2006	Nov 30,2011

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GOLDCORP CANADA LTD. Red Lake Gold Mines
Permit List
Original Date: November 2, 2006		Update:
Red Lake Complex
1.	PTTW # 0564-5TVP3K Mine workings and Balmer Lake	MOE	Dec. 22, 2003	March 31, 2008
2.	Certificate of Approval (Industrial Sewage) # 8559- 6RGQVW	MOE	July 18, 2006	N/A
3.	Certificate of Approval (Air) # 5339-6PXPP9	MOE	August 28, 2006	N/A
4.	Certificate of Approval (Waste) # A900417 generator # ON0344800	MOE		N/A
5.	Nuclear Permit # 11896-2-11-1	CNSC	September 2006	Nov 30,2011
Cochenour Complex
1.	Certificate of Approval (Industrial Sewage) # 9856- 6R5JFA	MOE	June 28, 2006	N/A
2.	Certificate of Approval (Air) # 4858-6N6MBG	MOE	April 18, 2006	N/A
3.	Waste Generator Number ON1674300	MOE		N/A
4.	Aggregate Permit # 14351	MNR	Current	N/A
	Balmer Complex
1.	Certificate of Approval (Air) # 2303-5VTQN2 (Shaft sinking)	MOE	Feb. 23, 2004	N/A

19.6.2Socio-economic and Environmental Conditions

RLGM mine site is located in the Red Lake District, which consists of 5 smaller communities of approximately 5,000 people. The complexes themselves are situated on the edges of the communities which make them a part of the community landscapes. Given these proximities operational and environmental considerations are elevated as are the company’s commitments to social, cultural, community support. As an example this can be demonstrated by RLGM’s continued support and operation of the Balmertown Community Recreation Facility.

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RLGM also participates with various local organizations such as the local municipal planning boards, hospital boards, economic development board as well as maintaining a community liaison committee.

19.6.3 Bonding, Reclamation and Closure

As per provincial and federal jurisdictions, Red Lake Gold Mines complies with the requisite bonding levels.

Three closure plans exist for the operations:  Cochenour Complex, Campbell Complex, Red Lake and Balmer Complexes combined. These documents have been created and or updated by Lorax Environmental and BZ Environmental.

Facility	Consultant	Date	MNDM
			Review/Status
Cochenour Wilanour	Lorax Environmental Inc	March 2005	Submitted – in process
Campbell Mine	BZ Environmental	April 2004	Submitted – Accepted
Red Lake Mine	Lorax Environmental Inc	March 2005	Submitted – in process

The closure plans outline the use of best available technology to decommission, reclaim and restore the mine site to a state that is as close to its pre-development condition as it is technically feasible. Currently RLGM is in the process of an internal review of the documents to ensure that all 3 plans reflect the most upto date scientific assessments, a standardized approach to issue management and to financial assurance.

Closure Plans

The closure plans reviewed three closure scenarios: temporary suspension, inactivity and final closure. In all cases, the plans presented satisfied the concerns of the Ontario government, First Nations and the general public.

Reclamation Plans

Reclamation activities are an ongoing process that run concurrently with production activities at the operations. Progressive reclamation initiatives include activities such as revegetating select areas or completing shaft/raise capping on sites that are inactive.

As sites continue thru the various phases of closure additional activities will undertaken which include but are not limited to:

•decommissioning of process plant and mine site

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•characterization studies

•demolition of site infrastructure

•seal mine access points

•mine site re-contour and re-vegetate

•tailings rip rap

The post closure environmental and long-term monitoring program is planned to last a minimum of ten years.

Waste Rock Management Plan

Waste Rock and ore are sampled for acid rock drainage (“ARD”) potential at all operations. The assessment program is currently under review by Lorax Environment to provide an integrated approach to a program that will encompass the entire business unit and verify the latest results.

Since there are no significant ARD issues related to the waste and ore from the Campbell and Red Lake Complexes, waste rock is routinely used as construction materials in site projects (i.e., roads, tails dams etc)  These two complexes each operate a waste rock stockpile which will be recontoured and revegetated post closure.

Tailings Management Plans

The Red Lake and Campbell Complexes each operate active tailings management areas (TMAs)  Each facility has been designed by engineering fims, Trow Engineering and AMEC engineering respectively. Annual geotechnical and facility inspections are conducted by these firms as well as additional engineering assessments and investigations to enhance tails storage strategies.

Currently, AMEC engineering is completely a review of both active tailings facilities and is developing individual Tailings Management Plans, updated waterbalances/storage reviews as well as an Operating, Maintenance and Surveillance manual.

Water treatment processes are in place or under construction at both facilities to address the destruction of cyanide and metals in solutions. All effluent discharges to the environment are in compliance with all applicable laws.

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Environmental Monitoring Programs

Red Lake Gold Mines has specific monitoring obligations that are required by the various Certificates of Approvals and legislative requirements including the Federal Metal Mining effluent regulations and the MISA regulations. The operating complexes have two unique systems which are currently being incorporated into an all inclusive document. These changes will not be completed until the 2007 when standard operating procedures, revised sampling schedules, and a single analytical laboratories will be finalized.

20.0CERTIFICATE OF QUALIFICATIONS AND CONSENT OF QUALIFIED PERSONS

CERTIFICATE OF QUALIFIED PERSON – Anthony Michael Stechishen

I, Anthony Michael Stechishen, P.Geo, as an author of this report entitled “Report on the Red Lake Gold Mines Property, Red Lake Mining Division, Northern Ontario (the “Technical Report”), prepared for Goldcorp Inc. and dated November 17, 2006, do hereby certify that:

1.I am a Senior Resource Geologist with Goldcorp Canada Ltd. at Red Lake Gold Mines, Balmertown, Ontario. My permanent residence is 276 Howey Street, P.O.Box 648, Red Lake, Ontario, P0V 2M0.

2.I graduated from the University Of Waterloo with a Bachelor of Science degree (Honours Applied Earth Sciences) in 1986.

3.I have worked as a geoscientist for a total of 20 years since my graduation from university. I have experience in mineral exploration, production geology, and relevant experience in resource estimation at the Red Lake Gold Mines - Campbell Complex.

4.I am a member in good standing of the Association of Professional Geoscientists of Ontario.

5.I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.I am currently an employee of Goldcorp at the Red Lake Gold Mines - Campbell Complex (formerly employed by Placer Dome (CLA) Limited at the Campbell Mine) and have been since 1999.

7.I am responsible for preparation of sections 12, 13, 14, 15, 16, and 18 of the Technical Report.

8.I am not independent of Goldcorp Inc. applying the test set out in Section 1.4 of NI 43-101.

9.I have been involved in regional exploration, underground production geology and resource estimation as an employee at the Campbell Complex prior to preparation of the Technical Report.

10.I have read NI 43-101, and the Technical Report has been prepared in compliance with NI 43-101 and Form 43-101F1.

11.To the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated 17^th day of November, 2006

/s/ Anthony Michael Stechishen

Anthony Michael Stechishen, P.Geo

Dean B. Crick

I, Dean B. Crick, M.Sc., P. Geo, as an author of this report entitled “Report on the Red Lake Gold Mines Property, Red Lake Mining Division, Northern Ontario” (the “Technical Report”), prepared for Goldcorp Inc. and dated November 17, 2006, do hereby certify that:

1.I am Manager of Mine Geology with Goldcorp Canada Ltd. at Red Lake Gold mines, P.O. Bag 2000, Balmertown, Ontario, P0V 1C0

2.I graduated with an Honours Bachelor of Science Degree from Brock University in 1986. In addition, I obtained a Masters of Science degree in Economic Geology from Laurentian University in 1991. I have fifteen years experience in exploration and mine production geology. I am a member in good standing with the Association of Professional Geoscientists of Ontario.

3.I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

4.I am currently employed at the Red Lake Gold mines and have been since December 2004.

5.I am responsible for preparation of sections 1, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, and 18

6.I am not independent of Goldcorp Inc. applying the test set out in Section 1.4 of NI 43-101.

7.My involvement in Red Lake Gold mines has included exploration and resource estimation
supervision in the roles of senior exploration geologist and subsequently manager of mine
geology.

8.I have read NI 43-101, and the Technical Report has been prepared in compliance with NI 43-101 and Form 43-101F1.

9.To the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated 17^th day of November, 2006

/s/ Dean B. Crick

Dean B. Crick, M.Sc., PGeo

Stéphane Blais

I, Stéphane Blais, Professional engineer, as an author of this report entitled “Report on the Red Lake Gold Mines Property, Red Lake Mining Division, Northern Ontario” (the “Technical Report”), prepared for Goldcorp Inc. and dated November 17, 2006, do hereby certify that:

1.I am Chief Engineer for Goldcorp Canada Ltd. at Red Lake Gold mines, P.O. Bag 2000, Balmertown, Ontario, P0V 1C0

2.I graduated with a Bachelor of Mining Eng. From the Université Laval in 1994.1 have twelve years experience in mining production and Engineering. I am a member in good standing with the Professional engineers of Ontario. I have practiced my profession continuously since 1995.

3.I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

4.I am currently employed at the Red Lake Gold mines and have been since July 2001.

5.I am responsible for preparation of sections 17,18 &19.

6.I am not independent of Goldcorp Inc. applying the test set out in Section 1.4 of NI 43-101.

7.My involvement in Red Lake Gold mines has included reserves estimation & supervision in the role of Chief Eng.

8.I have read NI 43-101, and the Technical Report has been prepared in compliance with NI 43-101 and Form 43-101F1.

9.To the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated 17^th day of November, 2006

/s/ Stéphane Blais

Stéphane Blais

Chief Eng. P.Eng.

CONSENT OF EXPERT
FILED BY SEDAR

November 17, 2006

British Columbia Securities Commission (Principal Regulator)

Ontario Securities Commission

Alberta Securities Commission

Saskatchewan Securities Commission

The Manitoba Securities Commission

Autorité des marchés financiers

Office of the Administrator, New Brunswick

Nova Scotia Securities Commission

Securities Commission of Newfoundland and Labrador

Registrar of Securities, Prince Edward Island

Registrar of Securities, Northwest Territories

Registrar of Securities, Government of the Yukon Territory

Registrar of Securities, Nunavut

Dear Sirs/Mesdames:

Re:Goldcorp Inc.

Report entitled “Report on the Red Lake Gold Mines Property, Red Lake Mining Division, Northern Ontario” dated November 17, 2006

Pursuant to Section 8.3 of National Instrument 43-101 Standards of Disclosure for Mineral Projects this letter is being filed as the consent of Dean Crick, P. Geo, Mine Geology Manager, Goldcorp Inc. to the public filing of the technical report entitled “Report on the Red Lake Gold Mines Property, Red Lake Mining Division, Northern Ontario” dated November 17, 2006.

Sincerely,

/s/ Dean Crick

Dean Crick, P. Geo

Mine Geology Manager

Goldcorp Inc.

CONSENT OF EXPERT
FILED BY SEDAR

November 17, 2006

British Columbia Securities Commission (Principal Regulator)

Ontario Securities Commission

Alberta Securities Commission

Saskatchewan Securities Commission

The Manitoba Securities Commission

Autorité des marchés financiers

Office of the Administrator, New Brunswick

Nova Scotia Securities Commission

Securities Commission of Newfoundland and Labrador

Registrar of Securities, Prince Edward Island

Registrar of Securities, Northwest Territories

Registrar of Securities, Government of the Yukon Territory

Registrar of Securities, Nunavut

Dear Sirs/Mesdames:

Re:Goldcorp Inc.

Report entitled “Report on the Red Lake Gold Mines Property, Red Lake Mining Division, Northern Ontario” dated November 17, 2006

Pursuant to Section 8.3 of National Instrument 43-101 Standards of Disclosure for Mineral Projects this letter is being filed as the consent of Anthony Stechishen, P. Geo, Senior Resource Geologist at Goldcorp Canada Ltd. to the public filing of the technical report entitled “Report on the Red Lake Gold Mines Property, Red Lake Mining Division, Northern Ontario” dated November 17,2006.

Sincerely,

/s/ Anthony Stechishen

Anthony Stechishen, P. Geo

Senior Resource Geologist

Goldcorp Canada Ltd.

CONSENT OF EXPERT
FILED BY SEDAR

November 17, 2006

British Columbia Securities Commission (Principal Regulator)

Ontario Securities Commission

Alberta Securities Commission

Saskatchewan Securities Commission

The Manitoba Securities Commission

Autorité des marchés financiers

Office of the Administrator, New Brunswick

Nova Scotia Securities Commission

Securities Commission of Newfoundland and Labrador

Registrar of Securities, Prince Edward Island

Registrar of Securities, Northwest Territories

Registrar of Securities, Government of the Yukon Territory

Registrar of Securities, Nunavut

Dear Sirs/Mesdames:

Re:Goldcorp Inc.

Report entitled “Report on the Red Lake Gold Mines Property, Red Lake Mining Division, Northern Ontario” dated November 17,2006

Pursuant to Section 8.3 of National Instrument 43-101 Standards of Disclosure for Mineral Projects this letter is being filed as the consent of Stephane Blais, P. Eng, Chief Engineer of Goldcorp Inc. to the public filing of the technical report entitled “Report on the Red Lake Gold Mines Property, Red Lake Mining Division, Northern Ontario” dated November 17,2006.

Sincerely,

/s/ Stephane Blais

Stephane Blais, P. Eng

Chief Engineer

Goldcorp Inc.

- 138 - November 17, 2006

21.0        REFERENCES

Andrews, A.J., Hugon, H., Durocher, M., Corfu, F. and Lavigne, M.J., (1986). The anatomy of the gold-bearing greenstoone belt: Red Lake, northwestern Ontario, Canada; in Proceedings of Gold ‘86, an International Symposium on the Geology of Gold Deposits, Konsult International Inc., Toronto, p3-22.

Atkinson, B.T. 1993b. Precambrian geology of Dome Township. Ontario Geological survey, Open File Report 5878, accompanied by OFM 231 scale 1:12000.

Atkinson, B.T. 1993c. Precambrian geology of the east part of Baird Township and HeysonTownship. Ontario Geological Survey, Open File Report 5870, 25p.

Atkinson, B.T. and Stone, D. 1993. Precambrian geology, Red Lake area. Ontario Geological Survey, Preliminary Map P.3227, scale 1:50000.

Cadieux, A-M., Dubé, B., Williamson, K., Malo, M. and Twomey, T. (2006). Charecterisation of hydrothermal alterations at the Red Lake mine, northwestern Ontario. Geological Survey of Canada, Current Research 2006-C2.

Chisholm, E.O. 1951. Geology of Balmer Township, Patricia portion of Kenora district (accompanied by map no. 1951-3). Ontario Department of Mines, vol. LX, pt. 10, 62p.

Damer, G.C. 1997. Metamorphism of hydrothermal alteration at the Red Lake Mine, Balmertown, Ontario. Unpublished M.Sc. thesis, Queen’s University, Kingston, Ontario. 195p.

Dube, B., Balmer, W., Sanborn-Barrie, M., Skulski, T., and Parker, J. 2000. A preliminary report on amphibolite-facies, disseminated-replacement-style mineralization at the Madsen gold mine, Red Lake, Ontario. Geological Survey of Canada, Current Research 2000-C17, 12p.

Dube, B., Williamson, K. and Malo, M. 2001a. Preliminary report on the geology and controlling parameters of the Goldcorp Inc. High Grade zone, Red Lake mine, Ontario. Geological Survey of Canada, Current Research 2001-C18.

Dube, B., Williamson, K. and Malo, M. 2001b. The Goldcorp high-grade zone, Red Lake mine, Ontario: a photographic atlas of the main geological features. Geological Survey of Canada, Open File 3890.

Dube, B., Williamson, K., and Malo, M. 2002. Geology of the Goldcorp Inc High Grade zone, Red Lake mine, Ontario: an update Geological Survey of Canada, Current Search 2002-C26. 23p.

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Dube, B., Williamson, K., and Malo, M. 2003. Gold mineralization within the Red Lake mine trend; example from the Cochenour-Willans mine area, RedcLake Ontario, with new key information from the Red Lake mine and potential analogy with the Timmins camp; Geological Survey of Canada, Current Research 2003-C21, 15 p.

Durocher, M.E., Burchell, P.S., Wallace, H., Andrews, A.J., Hugon, H and Atkinson, B.T. 1991. Geology and gold mineralization, Red Lake greenstone belt. Ontario Geological Survey, Preliminary Map P.3107, scale 1:50000.

Ferguson, S.A. 1966. Geology of Dome Township, district of Kenora. Ontario Department of Mines Geological Report 45, 98p. Accompanied by Map 2074, scale 1:12000.

Gammons, C.H. and Williams-Jones, A.E. 1997. Economic Geology, v 92, p45-59.

Heinrich, C.A. and Eadinton, P.J. 1986. Economic Geology, v 81, p511-527.

MacGeehan, P.J. and Hodgson, C.J., 1982. Environments of gold mineralization in the Campbell Red Lake land Dickenson mines Red Lake District, Ontario. In Geology of Canadian gold deposits. Edited by R.W. Hodder and W. Petruk. Canadian Institute of Mining and Metallurgy Special Volume 24, pp. 184-207.

MacGeehan, P.J. and Hodgson, C.J. 1981. Relationship of gold mineralization to volcanic and alteration features in the area of Campbell and Dickenson mines, Red Lake district. In Genesis of Archean Volcanic-Hosted Gold Deposits. Edited by E.G. Pye and R.G. Roberts. Ontario Geological Survey Miscellaneous Paper no. 97, pp. 94-110.

Mason, R. et al 2000. Proceedings of the 4th International Mining Geology Conference, Coolum, Queensland, AUSIMM/AIG, p 143-156.

Mathieson, N.A. 1982. Geology and mineralization in the area of the East South “C” orezone, Dickenson mine, Red Lake district, Northwestern Ontario. Unpublished M.Sc. thesis, Queen’s University, Kingston, Ontario. 155 p.

Mathieson, N.A. and Hodgson, C.J. 1984. Alteration, mineralization, and metamorphism in the area of the East South “C” ore zone, 24th level of the Dickenson mine, Red Lake, northwestern Ontario. Canadian Journal of Earth Sciences, vol. 21, pp. 35-52.

Matthai, S.K. et al 1995. Economic Geology, v 90, p 2123-2142. Ontario Geological Survey, 1951. Map 1951-3. Township of Balmer, scale 1:12000.

- 140 - November 17, 2006

Ontario Geological Survey, 1961. P47 Preliminary Geological and Geophysical Map and Township of Balmer, scale 1:12000. Revised 1961.

Ontario Geological Survey, 1961. P125. Preliminary Map. Dome Township, scale 1:12000. Ontario Geological Survey, 1965. Map 2074, Dome Township, scale 1:12000.

Ontario Geological Survey, 1978. Map p.1776A. Balmer Township Geological Series, scale 1:12000.

Parker, J.R. 2000. Gold mineralization and wall-rock alteration in the Red Lake greenstone belt: a regional perspective. In Summary of Field Work and Other Activities. Ontario Geological Survey, Open File Report 6032, p. 22-1-22-28.

Penczak, R.S. and Mason, R. 1999. Characteristics and origin of Archean premetamorphic hydrothermal alteration at the Campbell Mine, northwestern Ontario, Canada. Econ. Geol., Vol. 94, pp. 507-528.

Penczak, R.S. and Mason, R. 1997. Metamorphosed Archean epithermal Au-As-Sb-Zn-(Hg) vein mineralization at the Campbell Mine, northwestern Ontario. Econ. Geol., Vol. 92, pp. 696-719.

Pirie, J. 1977. Bateman-Balmer Township area, district of Kenora, Patricia portion. In Summary of field work, 1977, by the Geological Branch. Edited by V.G. Milne, O.L. White, R.B. Barlow and J.A. Robertson. Ontario Geological Survey Miscellaneous Paper no. 75, pp. 13-17.

Pirie, J. and Grant, A. 1978a. Bateman Township, district of Kenora (Patricia portion). Ontario Geological Survey, Preliminary Map P. 1569A, Geological Series, scale 1:12000.

Pirie, J. and Grant, A. 1978b. Balmer Township Area, district of Kenora (Patricia portion), Ontario Geological Survey Prelim. Map P.1976, Geological Series, scale 1:10560.

Rigg, D.M. 1980. Relationships between structure and gold mineralization in Campbell Red Lake and Dickenson mines, Red Lake District, Ontario. Unpublished M.Sc. thesis, Queen’s University, Kingston, Ontario. 153p.

Rigg, D.M. and Helmstaedt, H., 1981. Relations between structures and gold mineralization in the Campbell Red Lake and Dickenson Mines, Red Lake area, Ontario. In Genesis of Archean volcanic-hosted gold deposits. Edited by E.G. Pye and R.G. Roberts. Ontario Geological Survey, Miscellaneous Paper no. 97, pp. 111-127.

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Rigg, D.M., and Scherkus, E.W. 1983. Geology of the Wilmar mine, Red Lake Area. In The Geology of gold in Ontario. Edited by A.C. Colvine. Ontario Geological Survey, Miscellaneous Paper no. 110, pp. 159-173.

Rogers, J.A. 1992. The Arthur W. White mine, Red Lake area, Ontario: detailed structural interpretation the key to successful grade control and exploration. Canadian Institute of Mining and Metallurgy Bulletin, vol. 85, no. 957, pp 37-44.

Sanborn, M.M. 1984. Structural geology and mineralization of the West Carbonate Zone, Cochenour Willans gold mine, Red Lake. In Summary of field work, 1984, Ontario Geological Survey. Edited by J. Wood, O.L. White, R.B. Barlow, and A.C. Colvine. Ontario Geological Survey, Miscellaneous Paper no. 119, pp. 181-184.

Sanborn-Barrie, M., Skulski, T., Parker, J., and Dube, B. 2000. Integrated Regional analysis of the Red Lake greenstone belt and its mineral deposits, western Superior Province, Ontario. Geological Survey of Canada, Current Research 2000- C18, 16p.

Sanborn-Barrie, M., Skulski, T., and Parker, J. 2001. Three hundred million years of tectonic history recorded by the Red Lake greenstone belt, Ontario. Geological Survey of Canada, Current Research 2001-C19.

Sanborn-Barrie, M., Skulski, T., and Parker, J. 2004. Geology, Red Lake greenstone belt, western Superior Province, Ontario; Geological Survey of Canada, Open File 4594, scale 1:50000

Sketchley et al, 1997: Campbell Mine, Ontario, Ore Reserve Audit of November, 1997, Project Development Program for Placer Dome Canada Limited.

Stott, G.M. and Corfu, f., (1991). Geology of Ontario, OGS Special Volume 4, Part1. MNDM

Tarnocai, C.A., Hattori, K. and Cabri, L.J. 1997. Invisible gold in sulfides from the Campbell mine, Red Lake greenstone belt, Ontario: evidence for mineralization during peak of metamorphism: The Canadian Mineralogist, v.35, pt 4, p. 805-815.

Tarnocai, C.A. and Hattori, K. 1996. Int. Geol. Congress 30th, (Beijing) Abstr. p 765.

Tarnocai, C.A. and Hattori, K. 1994. Extended hydrothermal activity at the Campbell gold mine, eastern Red Lake greenstone belt, Uchi subprovince. In Geological Association of Canada Mineralogical Association of Canada annual meeting, program with abstracts, vol. XIX, p. A110.

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Tarnocai, C.A. and Hattori, K. and Stubens, T.C. 1998. Metamorphosed Archean Epithermal Au-As-Sb-Zn-(Hg) vein mineralization at the Campbell Mine, northwestern Ontario – A discussion, Econ. Geol., Vol. 93, pp 683-688

Twomey, T. and McGibbon, S. 2001. The geological setting and estimation of gold grade of the High-grade Zone, Red Lake Mine, Goldcorp Inc.; Exploration and Mining Geology, v. 10, p. 19-34.

Wall, V.J. 1987. Second Eastern Goldfields Geological Field Conference, Kalgoorlie. W.A., Geol. Soc. Australia, p 1-16.

Wall, V.J. et al 1995. In: Giant Ore Deposits – II, Controls on the Scale of Orogenic Magmatic-Hydrothermal Mineralization, Second Giant Ore Deposits Workshop, Kingston, Ontario, Canada, Alan H. Clark, Editor, Queen’s University, Kingston, Ontario, p 707-724.

Watts Griffith and McQuart Red Lake Mine December 2005 Audit. Internal document

Zhang, G., Hattori, K., and Cruden, A.R. 1997. Structural evolution of auriferous deformation zones at the Campbell mine, Red Lake greenstone belt, Superior Province of Canada: Precambrian Research, v. 84, pp. 83-103.

www.goldcorp.com; Company website, 2006

http://www.infomine.com/index/properties/RED_LAKE_MINE.html

http://www.infomine.com/index/properties/CAMPBELL_MINE.html

- 143 - November 17, 2006

22.0        ADDITIONAL FIGURES AND ILLUSTRATIONS

Figure 7.1

Figure 7.2

Figure 7.3

Figure 7.4

SGS Red Lake Duplicates	SGS Red Lake Duplicates
Gold Assays < 50 oz/t	Gold Assays < 5 oz/t

Figure 14.1        2005 Original assays versus Duplicate pulps for Red Lake Complex

SGS Red Lake Standards Results 2005 (Grav)

Figure 14.2 Red Lake Complex Standards Assay Results

Figure 18.7      Log normal histogram of selected Sulphide Zones.

Figue 18.8      Log normal probability plot for selected Sulphide Zones.

Figure 18.9: 10_00_F – F ZONE Probability plot uncapped 2ft (0.6m) composited chips	Figure 18.10: 10_00_F – F ZONE Probability plot uncapped 2ft (0.6m) composited chips

Figure 18.11: 10_00_F – F ZONE Histogram cap analysis for uncapped 2 ft (0.6m) composited chips	Figure 18.12: 10_00_F – F ZONE Histogram capping analysis for uncapped 2 ft (0.6m) composited drill holes

Figure 18.13: 308_27_56_561- 561 ZONE Probability plot uncapped 2 ft (0.6m) composited chips	Figure 18.14: 308_27_56_561- 561 ZONE Probability plot uncapped 2 ft (0.6m) composited drill holes

Figure 18.15: 308_27_56_561 – 561 ZONE Histogram for uncapped 2 ft (0.6m) composited chips	Figure 18.16: 308_27_56_561- 561 ZONE Histogram for uncapped 2 ft (0.6m) composited drill holes

Figure 18.17: 36_30_mm_tp – TP ZONE Probability plot uncapped 2 ft (0.6m) composited chips	Figure 18.18: 36_30_mm_tp Probability plot uncapped 2 ft(0.6m) composited drill holes

Figure 18.19: 36_30_mm_tp – TP ZONE Histogram for uncapped 2 ft (0.6m) composited chips	Figure 18.20: 36_30_mm_tp – TP ZONE Histogram for uncapped 2 ft (0.6m) composited drill holes

Figure 18.21: 42_36_MM_TP_SXC – TP ZONE Probability plot uncapped 2 ft (0.6m) composited chips	Figure 18.22: 42_36_MM_TP_SXC – TP ZONE Probability plot uncapped 2 ft (0.6m)composited drill holes

Figure 18.23: 42_36_MM_TP_SXC – TP ZONEHistogram for uncapped 2 ft (0.6m) composited chips	Figure 18.24: 42_36_MM_TP_SXC - TP ZONE Histogram for uncapped 2 ft (0.6m) composited drill holes

Figure 18.25: 42_36_MM_TP_SXC – TP ZONE Probability plot uncapped 2 ft (0.6m) composited chips	Figure 18.26: 42_36_MM_TP_SXC – TP ZONE Probability plot uncapped 2 ft (0.6m) composited drill holes

Figure 18.27: 42_36_MM_TP_SXC – TP ZONE Histogram for uncapped 2 ft (0.6m) composited chips	Figure 18.28: 42_36_MM_TP_SXC – TP ZONE Histogram for uncapped 2 ft (0.6m) composited drill holes

Figure 18.29 42_36_DC – DC ZONE Probability plot for 2 ft (0.6m) uncapped composited chips	Figure 18.30 42_36_DC – DC ZONE Probabily plot for uncapped 2 ft (0.6m) composited drill holes

Figure 18.31 42_36 DC – DC ZONE Histogram for uncapped for 2 ft (0.6m) uncapped composited chips	Figure 18.32 42_36_DC – DC ZONE Histogram for uncapped for uncapped 2 ft (0.6m) composited drill holes

APPENDIX A
LAND CLAIM STATUS

- A1 - November 17, 2006

Table 5.1List of Claims registered for the Red Lake Gold Mines

Company	Mining Division	Land Registry Office	Township	Tenure Number		# Units	Hectares	Tenure Type
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19689	1	7.77	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19690	1	10.522	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19691	1	9.712	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19692	1	12.586	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19696	1	17.725	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19697	1	21.732	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19698	1	17.847	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19699	1	16.875	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19700	1	9.591	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19701	1	15.783	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19702	1	11.331	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20141	1	19.304	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20142	1	7.39	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20143	1	10.445	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	23105	1	7.062	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19493	1	13.929	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19494	1	21.375	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19495	1	16.77	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19496	1	3.313	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19496	1	0.089	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19496	1	15.249	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19496	1	0.038	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19496	1	0.083	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19497	1	23.055	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19497	1	0.08	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19498	1	12.137	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19499	1	0.038	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19499	1	0.046	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19499	1	15.512	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19499	1	0.046	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19499	1	0.089	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19499	1	0.089	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19499	1	0.038	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19499	1	7.286	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19500	1	18.365	Patented MR & SR

- A2 - November 17, 2006

Company	Mining Division	Land Registry Office	Township	Tenure Number		# Units	Hectares	Tenure Type
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19500	1	0.038	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19501	1	14.977	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19502	1	18.001	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19502	1	14.01	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19503	1	17.685	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19503	1	0.089	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19504	1	18.454	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19505	1	33.832	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19506	1	23.188	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19507	1	21.125	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19508	1	12.829	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19509	1	16.956	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19510	1	16.349	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19511	1	15.014	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19512	1	15.095	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19513	1	11.845	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19514	1	14.083	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20779	1	16.859	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20780	1	12.165	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20781	1	21.23	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20782	1	16.993	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20783	1	21.295	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20793	1	21.537	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20794	1	24.196	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20795	1	25.082	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20796	1	25.366	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20797	1	19.898	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20798	1	27.219	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	20799	1	19.069	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19643	1	7.535	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19644	1	26.305	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19645	1	14.164	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19646	1	12.222	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19647	1	12.829	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19973	1	16.592	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19974	1	15.459	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19975	1	12.141	Patented MR & SR

- A3 - November 17, 2006

Company	Mining Division	Land Registry Office	Township	Tenure Number		# Units	Hectares	Tenure Type
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19976	1	15.783	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19977	1	7.365	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19978	1	12.262	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19979	1	14.002	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19980	1	21.165	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	19981	1	21.893	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	21286	1	25.091	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	21287	1	17.806	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	21288	1	13.112	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	21289	1	11.129	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	21290	1	13.921	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	21291	1	19.983	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	21292	1	11.21	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	21237	1	28.692	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	21238	1	20.72	Patented MR & SR
Goldcorp Inc.	Red Lake	Kenora	Balmer	KRL	21239	1	24.767	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	CLM	304	1	0.72	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	10867-LO	1	14.75	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	10871-LO	1	6.3	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	10919-LO	1	11.77	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	11122-LO	1	4.15	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	11123-LO	1	6.13	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	11124-LO	1	5.67	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19666	1	15.76	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19667	1	18.54	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19668	1	11.43	Patented MR & SR

- A4 - November 17, 2006

Company	Mining Division	Land Registry Office	Township	Tenure Number		# Units	Hectares	Tenure Type
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19669	1	18.78	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19670	1	18.07	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19671	1	10.57	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19721	1	11.95	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19722	1	25.71	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19723	1	16.65	Patented MR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19724	1	25.88	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19725	1	18.19	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	19726	1	6.99	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20068	1	22.6	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20069	1	22.28	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20070	1	10.79	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20071	1	16.09	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20072	1	10.87	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20073	1	20.63	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20074	1	20.33	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20075	1	19.89	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20076	1	16.33	Patented MR

- A5 - November 17, 2006

Company	Mining Division	Land Registry Office	Township	Tenure Number		# Units	Hectares	Tenure Type
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20252	1	11.73	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20253	1	29.72	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20254	1	23.55	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20255	1	12.49	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20256	1	22.09	Patented MR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20345	1	19.02	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20346	1	18.5	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20459	1	13.48	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20460	1	18.77	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20461	1	17.07	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20462	1	11.34	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20603	1	11.7	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20604	1	8.43	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20605-LO	1	15.88	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20606-LO	1	16.6	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20607-LO	1	12.74	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20731	1	0.73	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20755	1	4.21	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20824-LO	1	11.91	Lic. of Occupation

- A6 - November 17, 2006

Company	Mining Division	Land Registry Office	Township	Tenure Number		# Units	Hectares	Tenure Type
Canada Ltd.
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20825-LO	1	13.55	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20826-LO	1	5.01	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20829-LO	1	20.19	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20830-LO	1	16.99	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20831-LO	1	13.55	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	21953	1	9.03	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	21954	1	16.86	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	23059	1	1.02	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	27179	1	4.07	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	17956	1	16.19	Patented MR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	17957	1	18.86	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	17958	1	12.99	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	17959	1	16.72	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	18327	1	9.46	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	18328	1	13.35	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	18329	1	17.07	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	18330	1	15.5	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	18331	1	14.69	Patented MR & SR

- A7 - November 17, 2006

Company	Mining Division	Land Registry Office	Township	Tenure Number		# Units	Hectares	Tenure Type
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	18332	1	8.51	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	18333	1	12.23	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	18334	1	14.95	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	18335	1	12.16	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Dome	KRL	18336	1	17.48	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	10805-LO	1	3.39	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	10806-LO	1	6.62	Lic. of Occupation
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20840	1	14.29	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20841	1	25.03	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20842	1	25.1	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer	KRL	20843	1	14	Patented MR & SR
Goldcorp Canada Ltd.	Red Lake	Kenora	Balmer/Do me	KRL	21961	1	7.77	Patented MR & SR