Technical Report on Fosterville Gold Mine, Victoria, Australia.

Prepared for Northgate Minerals Corporation

March 25, 2008

Simon Hitchman	Ian Holland	Brad Evans
MAusIMM	MAusIMM	MAusIMM
Northgate Minerals	Northgate Minerals	Mining Plus
McCormicks Road	McCormicks Road	PO Box 205
Fosterville, VIC, 3557	Fosterville, VIC, 3557	Cloverdale, WA, 6985

Table of Contents
1.	Summary	7
1.1	Location	7
1.2	History and Ownership	7
1.3	Geology and Mineralisation	8
1.4	Mineral Resources and Mineral Reserves	9
1.5	Current Status	11
1.6	Conclusions and Recommendations	11
2.	Introduction	13
2.1	Terms of Reference	13
2.2	Sources of Information	13
2.3	Field Involvement of Qualified Persons	13
2.4	Definitions	14
2.5	Grids	16
3.	Reliance on Other Experts	16
4.	Property Description and Location	17
5.	Accessibility, Climate, Local Resources, Infrastructure and Physiography	20
6.	History	21
7.	Geological Setting	22
7.1	Regional Geology	22
7.2	Deposit Geology	24
8.	Deposit Types	26
9.	Mineralisation	27
9.1	Controls on Primary Mineralisation	28
9.1.1	Fosterville Fault Zone	28
9.1.1.1	Central and Hunts Areas	28
9.1.1.2	Wirrawilla – Daley’s Hill Area	30
9.1.2	Robbins Hill Area	31
9.2	Controls on Oxide Mineralisation	31
10.	Exploration	32
10.1	Pre-1992	32
10.2	1992-2001	32
10.3	2001-Present	33
11.	Drilling	35
11.1	Pre-1992	35
11.2	1992-2001	35
11.3	Current Drilling Methods	36
12.	Sampling Method and Approach	40
13.	Sample Preparation, Analyses and Security	43
13.1	Elements	43
13.2	Description of Analytical Techniques	44
13.3	Sample and Data Security	45
13.3.1	Sample Security	45
13.3.2	Data Security	45
13.4	Adequacy of Procedures	46
14.	Data Verification	46
14.1	Database Validation	46
14.2	QAQC	47
14.2.1	Assays	47

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14.2.2	Surveys	48
14.3	Data Verification	48
15.	Adjacent Properties	49
16.	Mineral Processing and Metallurgical Testing	49
16.1	Results	51
17.	Mineral Resource and Mineral Reserve Estimates	55
17.1	Central Area	59
17.1.1	Geological Models	60
17.1.2	Domaining	61
17.1.2.1	Falcon Domain	62
17.1.2.2	Phoenix Domain	62
17.1.2.3	Ellesmere Domain	62
17.1.2.4	Harrier & Splays Domain	63
17.1.3	Drilling Data	63
17.1.3.1	Compositing	64
17.1.3.2	Variography	66
17.1.4	Resource Modelling	67
17.1.4.1	Block Models	67
17.1.4.2	Search Criteria	68
17.1.4.3	Bulk Density	70
17.1.5	Mineral Resource Classification	71
17.1.6	Mineral Reserve Estimate	71
17.1.6.1	Mineral Reserve Block Design	72
17.1.6.2	Open Stope Design and Reserve Parameters	72
17.1.6.3	Modified Longitudinal Sublevel Caving Design and Reserve Parameters	75
17.1.6.4	Cut-off Grades	76
17.1.7	Depletion and Results	77
17.2	Northern Area	78
17.2.1	Area Geology	78
17.2.2	Geological Models	80
17.2.3	Domaining	81
17.2.4	Drilling Data	82
17.2.4.1	Compositing and Coding	82
17.2.4.2	Variography	83
17.2.5	Resource Modelling	83
17.2.5.1	Block Models	83
17.2.5.2	Search Criteria	83
17.2.5.3	Bulk Density	84
17.2.6	Mineral Resource Classification	84
17.2.7	Results	85
17.3	Southern Area	85
17.3.1	Area Geology	86
17.3.2	Geological Models	87
17.3.3	Domaining	87
17.3.3.1	Fosterville Domain	88
17.3.3.2	Harrier Splays Domain	88
17.3.3.3	Daley’s Hill Domains	89
17.3.3.4	Material Domains	89
17.3.4	Drilling Data	90
17.3.4.1	Compositing and Coding	90
17.3.4.2	Variography	90

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17.3.5	Resource Modelling	91
17.3.5.1	Block Models	91
17.3.5.2	Search Criteria	91
17.3.5.3	Bulk Density	92
17.3.6	Mineral Resource Classification	92
17.3.7	Results	92
17.4	Robbins Hill Area	93
17.4.1	Area Geology	93
17.4.1.1	O’Dwyer's to Robbins Hill Area – Geological Overview	93
17.4.1.2	Farley’s-Sharkey’s Area - Geological Overview	95
17.4.2	Geological Models	99
17.4.3	Domaining	99
17.4.3.1	Gold Domains	100
17.4.3.2	Oxidation Domains	100
17.4.4	Drilling Data	101
17.4.4.1	Compositing and Coding	101
17.4.4.2	Variography	102
17.4.5	Resource Modelling	102
17.4.5.1	Block Models	102
17.4.5.2	Search Criteria	103
17.4.5.3	Bulk Density	104
17.4.6	Mineral Resource Classification	104
17.4.7	Results	104
18.	Other Relevant Data and Information	105
19.	Interpretation and Conclusions	105
20.	Recommendations	106
21.	References	107
22.	Additional Requirements for Technical Reports on Development Properties and Production Properties	110
22.1	Mining Operations	110
22.2	Recoverability	112
22.3	Markets	112
22.4	Contracts	112
22.5	Environmental Considerations	113
22.6	Taxes	113
22.7	Capital and Operating Estimates	115
22.8	Economic Analysis	115
22.9	Payback	117
22.10	Mine Life	117
23.	Date and Signature Page	118
24.	Certificates of Qualified Persons	119

List of Illustrations
Figure 1: Location map.	7
Figure 2: Plan showing Fosterville mining lease, surrounding exploration leases and royalty coverage.	18
Figure 3: Schematic cross-section (EW) through the upper crust in the western Lachlan Fold Belt.	23
Figure 4: Cross-section though Phoenix and Ellesmere Shoots showing mineralization domains, faults, bedding formlines and the location of hanging wall black shale.	25

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Figure 5: Schematic cross section showing the relationship between mineralisation, faulting and bedding orientation.	30
Figure 6: Underground diamond drilling traces (green) in footwall of Phoenix and Falcon ore zones.	39
Figure 7: Face sample field duplicate results	42
Figure 8: Longitudinal projection showing location of the metallurgical testwork conducted from 1996-2003.	50
Figure 9: Plan showing mining lease and the area covered by each of the block models.	58
Figure 10: Cross-section though Phoenix and Ellesmere Shoots showing mineralization domains, faults, bedding formlines and the location of hanging wall black shale.	59
Figure 11: Section showing elements used in the construction of mineralisation domain wireframes.	60
Figure 12: Long projection of mineralisation domains in the Central Area	63
Figure 13: Downhole compositing where the width of the domain object is not a multiple of two.	65
Figure 14: Search ellipses for Domains 1 (Falcon) and 2 (Phoenix).	69
Figure 15: Section through the Falcon Block Model (8600mN), showing the Phoenix and Ellesmere mineralisation envelopes, and ore and waste grades.	69
Figure 16: Longitundinal section showing resource classification for Falcon and Phoenix ore zones (ore drives and stope shapes only) - Measured (blue), Indicated (green) and Inferred (red).	71
Figure 17: An example of an open stope reserve wireframe design	73
Figure 18: Longitundinal projection displaying the mining blocks referred to in Table 18 above.	75
Figure 19: Cross section 10,900mN through Hunt’s pit showing the relationship between the Hunt’s Fault, bedding and the set of splays that strike obliquely to the fault.	80
Figure 20: Schematic Wirrawilla geological cross section showing folded bedding, anticline and syncline axial planes, faulting and mineralisation	86
Figure 21: Longitudinal projection, looking west at Southern Model area. Shown are mineralisation wireframes, drill traces and open pits (yellow).	88
Figure 22: Plan view of ODW-RH area, showing oxide pits and mineralisation. In green are RH and ODW sub-vertical porphyry dykes. West dipping faults are in blue, browns, purples and magenta.	95
Figure 23: Cross-section on 12300mN	96
Figure 24: Cross-section on 12400mN	97
Figure 25: Cross-section on 12650mN	98
Figure 26: Longitudinal section showing actual and proposed mining layout as at November 2007.	110
Figure 27: Schematic Ore Treatment Flowsheet	111
Figure 28: Sensitivity analysis.	117
List of Tables
Table 1: Mineral Resources for Fosterville Gold Mine as at December 31, 2007.	9
Table 2: Mineral Reserves for Fosterville Gold Mine as at December 31, 2007.	10
Table 3: Grid coversion reference points	16
Table 4: Climate statistics from the Bendigo Airport weather station, 1991-2007 (Bureau of Meteorology, 2008).	20
Table 5: Mined production data for Fosterville for the period 2004-2007.	22
Table 6: Drill hole prefixes for all drilling on the Fosterville Fault Zone between 6000mN and 10000mN.	40

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Table 7: Analysed elements by method and time period.	43
Table 8: Details of the metallurgical samples collected during 1996-2003 and the testwork undertaken	51
Table 9: Flotation test results	52
Table 10: Summary of samples used for BIOX® and cyanide leach testwork.	53
Table 11: Summary of BIOX® variability testwork.	53
Table 12: Summary of cyanide leach of BIOX® variability residues.	54
Table 13: Mineral Resources for Fosterville Gold Mine as at December 31, 2007.	55
Table 14: Mineral Reserves for Fosterville Gold Mine as at December 31, 2007.	56
Table 15: Composite statistics by composite length.	65
Table 16: Central Area block model dimensions.	67
Table 17: Bulk density values used in the resource model.	70
Table 18: Recovery and dilution factors for the reserve blocks displayed in Figure 18 below.	74
Table 19: Cut-off grade assumptions used, AUD$ denomination.	77
Table 20: Proposed drilling programs for 2008	106
Table 21: Production forecast for the years 2008-2014 (LOM plan).	112
Table 22: Capital and operating cost estimates from the LOM plan completed in August 2007.	115
Table 23: Cash flow summary for Fosterville for the period 2008-2014.	116

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1. Summary

1.1 Location

The Fosterville Gold Mine is located approximately 20 kilometres north-east of Bendigo and 130 kilometres north of Melbourne in Victoria, Australia (Figure 1).

1.2 History and Ownership

Gold was first discovered in the Fosterville area in 1894 with activity continuing until 1903 for a total of 28,000 ounces of production. Mining in this era was confined to the near-surface oxide material. Aside from a minor tailings retreatment in the 1930’s, activity resumed in 1988 with a further tailings retreatment program conducted by Bendigo Gold Associates which ceased in 1989. Mining recommenced in 1991 when initially Brunswick and then Perseverance from 1992 commenced heap leaching operations from shallow oxide open pits. Between 1988 and the cessation of oxide mining in 2001, a total of 240,000 ounces of gold were poured (Roberts et al, 2003).

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A feasibility study into a sulphide operation was completed by Perseverance in 2003 with construction and open pit mining commencing in early 2004. Commercial production commenced in April 2005 and up to the end of December 2007 had produced 263,900 ounces of gold. In October 2007, Perseverance announced that it had entered into an agreement with Northgate Minerals Corporation (Northgate) to acquire the company. Full control passed to Northgate in February 2008.

1.3 Geology and Mineralisation

The Fosterville Goldfield is located within the Bendigo Zone in the Lachlan Fold Belt. The deposit is hosted by an interbedded turbidite sequence of sandstones, siltstones and shales. This sequence has been metamorphosed to sub-greenschist facies and folded into a set of upright, open to closed folds. The folding resulted in the formation of a series of bedding parallel laminated quartz veins. Although visually similar to their mineralised equivalents at Bendigo (20 kilometres away), these laminated quartz veins at Fosterville are effectively unmineralised.

Mineralisation at Fosterville is controlled by late brittle faulting. These late brittle faults are generally steeply west dipping reverse faults with a series of moderately west dipping reverse splay faults formed in the footwall of the main fault. Primary gold mineralisation occurs as disseminated arsenopyrite and pyrite forming as a selvage to veins in a quartz – carbonate veinlet stockwork. The mineralisation is structurally controlled with high grade zones localised by the geometric relationship between bedding and faulting. Mineralised shoots are typically 4 to 15 metres thick, 50 to 150 metres up / down dip and 300 to 1500 metres+ down plunge.

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1.4 Mineral Resources and Mineral Reserves

Table 1 below outlines the Mineral Resources for the Fosterville Gold Mine as at December 31, 2007.

Mineral Resources - Fosterville
			Measured			Indicated			Inferred
Classification		Tonnes(kt)	Grade(g/t Au)	Insitu Gold (kOz)	Tonnes(kt)	Grade(g/t Au)	Insitu Gold (kOz)	Tonnes(kt)	Grade(g/t Au)	Insitu Gold (kOz)
Fosterville Fault Zone Sulphide Resources
	Upper	2,326	2.45	183	404	1.87	24	146	2.23	11
Central Area
	Lower	703	4.71	107	5,766	4.74	879	6,108	3.58	703
Southern Area
	Upper	152	2.04	10	507	2.32	38	1,166	1.93	72
	Lower							5,648	3.34	607
Northern Area	Upper	333	1.83	20	240	1.23	10	25	0.99	1
	Lower
Robbin's Hill Area Sulphide Resources
Combined	Upper	39	1.19	1	2,303	1.31	97	1,398	1.44	65
	Lower				296	2.59	25	853	2.53	69
Sulphide Upper		2,849	2.34	214	3,454	1.52	169	2,735	1.69	148
Sulphide Lower		703	4.71	107	6,062	4.64	903	12,609	3.40	1,380
Total Sulphide		3,553	2.81	321	9,516	3.50	1,072	15,344	3.10	1,529
Total Oxide		606	1.17	23	1,857	1.10	66	598	1.36	26
Total Oxide & Sulphide		4,159	2.57	343	11,372	3.11	1,138	15,942	3.03	1,555

Table 1: Mineral Resources for Fosterville Gold Mine as at December 31, 2007.

Notes:

1.

For the Mineral Resource estimate, the Qualified Person for the Central Area is Ian Holland and for all other areas is Simon Hitchman. Their details and qualifications can be seen in Section 23 of this report.

2.

The Mineral Resources for the Central Area reported in Table 1 are inclusive of the Mineral Reserves for the same area (reported in Table 2 below).

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3.

Cut-off grades applied to Mineral Resources are 0.5 g/t Au and 0.7 g/t Au for oxide and sulphide mineralisation respectively above 5050mRL (approximately 100 metres below surface) which is deemed to be potentially open-pittable. The Mineral Resource cut-off grade applied for material below this point is 2.0 g/t Au.

4.

Mineral Resources have been rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summation may occur due to rounding.

Table 2 below outlines the Mineral Reserves for the Fosterville Gold Mine as at December 31, 2007.

Mineral Reserves - Fosterville
	Proven			Probable			Total
Classification	Tonnes (kt)	Grade (g/t Au)	Contained Gold (kOz)	Tonnes (kt)	Grade (g/t Au)	Contained Gold (kOz)	Tonnes (kt)	Grade (g/t Au)	Contained Gold (kOz)
Central Area
Phoenix	49	3.45	5	4,444	4.48	640	4,493	4.47	646
Falcon	87	4.00	11	95	3.69	11	182	3.84	22
Ellesmere				550	4.61	82	550	4.61	82
Kink				390	3.79	48	390	3.79	48
Total	136	3.80	17	5,480	4.43	781	5,616	4.41	798

Table 2: Mineral Reserves for Fosterville Gold Mine as at December 31, 2007.

Notes:

1.

For the Mineral Reserves estimate, the Qualified Person is Brad Evans. His details and qualifications can be seen in Section 23 of this report.

2.

The Mineral Reserve estimate used a gold price of AUD$750/ounce. The cut-off grades applied ranged from 1.2 g/t to 3.6 g/t Au for underground sulphide ore depending upon width, mining method and ground conditions.

3.

Dilution of 5-30% and mining recovery of 70-95% were applied to the Mineral Reserves dependant upon mining method.

4.

Mineral Reserves have been rounded to 1,000 tonnes, 0.1 g/t Au and 1,000 ounces. Minor discrepancies in summation may occur due to rounding.

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The Mineral Resources and Mineral Reserves reported above are largely the result of work undertaken in June 2007 and reported by Perseverance under Australian reporting requirements in accordance with the JORC Code. Although there have been some minor updates, the majority of the models remain unchanged from this period aside from being depleted for material mined from July 1, 2007 to December 31, 2007. Where there has been new data acquired (drilling or otherwise) in this period, it is the opinion of the authors that this new data has not made a material difference to this estimate.

In all cases, the Qualified Persons have reconciled the estimates to CIM standards as

prescribed by NI 43-101.

1.5 Current Status

Since the commencement of commercial gold production in April 2005, the sulphide plant at Fosterville Gold Mine has produced 263,900 ounces of gold up to the end of December 2007. This production was initially sourced solely from open cut mining with underground mining starting to contribute in late 2006. With the Harrier open cut being completed in December 2007 and no current plans for further open cut mining, the underground operations are now the sole source of ore for the plant.

1.6 Conclusions and Recommendations

The authors have made the following interpretations and conclusions:

1.

The understanding of the geological controls on mineralisation at Fosterville is high. Primary mineralisation is structurally controlled with high grade zones localised by the geometric relationship between bedding and faulting. This predictive model has lead to considerable exploration success in following the down-plunge extensions of high grade mineralisation.

2.

By the same token, this understanding has lead to the development of robust geological and resource models underpinning the Mineral Resource and Mineral Reserve estimates. The relationship between mineralisation and the controlling structural/stratigraphic architecture means that quality geological interpretation is critical to producing quality resource/reserve estimates.

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3.

The modifying factors used to convert the Mineral Resources to Mineral Reserves have been refined with the operating experience gained since underground production commenced in September 2006. In particular, the robustness of the mining recovery and dilution estimates has improved with experience relative to the pre-mining assessments.

The following recommendations are made:

1.

Further mine lease exploration/resource development activities should be pursued. Given the strong understanding of geological controls on mineralisation, this would be considered likely to yield additional resources and reserves. Particular areas that are recommended to focus upon are the Wirrawilla zone (5300-6100mN) in the Southern Area and the down-plunge extension of the Phoenix orebody beyond the current southern terminus of Mineral Reserves at approximately 7400mN. Detailed programs including costing have not yet been prepared.

2.

 The infill drilling program should be continued with an aim to maintain at least 12 months of reserves drilled out to 25 metre centres (or closer where necessary). Given the south plunging geometry of the Phoenix orebody, this is being conducted from a dedicated development drive in the hangingwall of the Fosterville Fault (western side). The current infill drilling budget for 2008 includes 30,000 metres of drilling at an estimated cost of approximately A$3.6 million. As the decline and mining front continues to move south and lower, further hangingwall drives will be required to be developed. This work and the associated drilling have not been costed in detail.

3.

With this additional drilling data and further ongoing operational experience, it is recommended that mining recovery and dilution factors are reviewed and refined on an ongoing basis.

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2. Introduction

2.1 Terms of Reference

The Fosterville Gold Mine has recently changed ownership through the acquisition of Perseverance by Northgate (transaction completed in February 2008). Prior to this transaction, Perseverance as an Australian-listed company reported resources and reserves pursuant to the JORC Code. The latest update was as at June 30, 2007. As a Canadian listed company, Northgate are required to report resources and reserves pursuant to National Instrument 43-101. This report has been prepared for Northgate to reflect the resource/reserve position for the Fosterville Gold Mine as at December 31, 2007 and to reconcile the JORC Code estimates to CIM standards as required by NI 43-101.

2.2 Sources of Information

Information and data for this report have been obtained from a range of sources including the personal work of the authors, contributions by other Fosterville personnel and reports from a range of external consultants. Where direct reference has been made to either a public or internal report, this reference has been cited and details can be found in Section 21.

2.3 Field Involvement of Qualified Persons

Simon Hitchman is currently employed as the Exploration Manager (Acting) for Northgate, based at the Fosterville Gold Mine. He has worked on site at Fosterville since July 2004 in exploration and resource definition including responsibility for resource estimation.

Ian Holland is currently employed as the Principal Mine Geologist for Northgate, based at the Fosterville Gold Mine. He has worked on site at Fosterville has worked on site at Fosterville since July 2007 with responsibility for all aspects of mining geology including resource and reserve estimation.

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Brad Evans was previously employed as the Senior Mining Engineer for Fosterville for the period September 2006 to January 2008 with responsibility for long term mine planning, including reserve estimation. Since January 2008, he has been employed by Mining Plus Pty Ltd as a Consulting Engineer but has continued to be based at Fosterville.

2.4 Definitions

Au	gold
As	arsenic
S	sulphur
Sb	antimony
NCC	non-carbonate carbon
AUD	Australian Dollars
g/t	grams per (metric) tonne
oz	ounce (31.104 grams)
m	metre
km	kilometre
t	(metric) tonne (2204.6 lb or 1.1023 short tons)
BCM	Bank (in situ) cubic metre
MGA94	Map grid of Australia 1994
AMG66	Australian Map Grid 1966
AMG84	Australian Map Grid 1984
AHD	Australian Height Datum (mean sea level)
RL	Reduced level (elevation)
E	Easting
N	Northing
FGP	Fosterville Gold Project
Perseverance	Perseverance Corporation Ltd., a wholly owned subsidiary of Northgate Minerals Corporation
Northgate	Northgate Minerals Corporation
BGA	Bendigo Gold Associates Ltd., owner of the FGP prior to Brunswick
Brunswick	Brunswick Mining N.L.., owner of the FGP prior to Perseverance

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BIOX	Proprietary bacterial oxidation technology licensed from Goldfields Ltd.
kg/m3	kilograms per cubic metre (unit of density)
ha	hectare (2.4711 acres)
CV	coefficient of variation
IDW	inverse distance weighting
OK	ordinary kriging
CIL	carbon in leach
DD	Diamond drill hole
RC	Reverse circulation drill hole
RAB	Rotary air blast drill hole
GC	Grade control
JORC Code	Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves (prepared by the Joint Ore Reserves Committee).
ML	Mining Lease
tpa	tonnes per annum
SAG	semi-autogenous grinding
CCD	counter current decantation
PQ	85.0 mm diameter diamond drill core
HQ	63.5 mm diameter diamond drill core
HQ2	61.1 mm diameter diamond drill core
NQ	47.6 mm diameter diamond drill core
NQ2	45.1 mm diameter diamond drill core
LTK60	49.0 mm diameter diamond drill core
Spear sampling using a tube (‘spear’) to collect a sample for assay from a sample bag of RC or RAB drill chips (this method is not equi-probable as it is susceptible to density segregation in the sample bag)
Riffle splitter	a device comprising tiers of ‘riffles’ for equi-probable splitting of dry particulate matter (e.g. drill chips), each tier yields a 50:50 split.
ROM	run of mine

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2.5Grids

Note that all eastings, northings, elevations (RL) and azimuths in the text reference to the local Fosterville Mine grid. The Fosterville Mine grid is a plane affine grid and can be referenced to AGD66 using the two reference points contained in Table 3 and -5000mRL (AHD). Fosterville Mine grid north is 13°20’ west from true north and 21° west from magnetic north.

Point 1: MIN5404 Lease peg SE of Daley's Hill
Coordinate System	N	E
AMG Zone 55 (AGD66)	5930654.500	277898.300
Fosterville Mine Grid	4786.030	2177.630
Point 2: MIN5404 Lease peg at NE corner
Coordinate System	N	E
AMG Zone 55 (AGD66)	5938863.100	278294.900
Fosterville Mine Grid	12713.150	4343.140

Table 3: Grid coversion reference points.

3. Reliance on Other Experts

The authors have prepared this report from a range of sources including their personal work, contributions from other Fosterville personnel and reports from a range of external consultants. Where input has been received from these sources, all attempts have been made by the authors to review and verify the contained assumptions and conclusions.

There are areas that were outside of the expertise of the authors, including environmental, legal, taxation, title and permitting considerations. For these areas, the authors are reliant upon the work undertaken by Allens Arthur Robinson and KPMG on behalf of Northgate during the due diligence in relation to the Perseverance acquisition.

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4. Property Description and Location

The Fosterville Gold Mine is located about 20 kilometres north east of Bendigo and 130 kilometres north of Melbourne in Victoria, Australia (see Figure 1).

The Fosterville Gold Mine and all associated infrastructure including the tailings dam and waste dumps are located on Mining Lease 5404 (Figure 2), which is 100% owned by Perseverance Exploration Pty Ltd. MIN5404 was initially granted as ML1868 on 24th August 1990. The licence later merged with adjoining lease MIN4877, resulting in MIN5404, which has a total area 17.03 km², and is active until the 24^th August, 2020. MIN5404 is located at centroid coordinates 276,599.72E and 5935,134.9N using Map Grid of Australia Zone 55 (GDA94) coordinate projection (or 144° 29’ 56.9" Longitude and 36 ° 42’ 11.6" Latitude).

17

Figure 2: Plan showing Fosterville mining lease, surrounding exploration leases and royalty coverage.

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The boundaries of land covered by the mining licence are accurately surveyed and marked on the ground with posts, trenches and information plates in accordance with the Mineral Resources Development Regulations 2002.

Northgate also holds title through Perseverance of six surrounding exploration licences totalling 1,437 km2. These exploration licences encompass the entire known strike extent of the Fosterville Goldfield. In Victoria, exploration licences are renewable annually subject to adequate exploration expenditure (Figure 2).

Within MIN5404, there is a 2.5% gold royalty payable to New Holland Mining Ltd, now Nu Energy Capital Limited for the area outlined by MIN4877 in the north eastern portion of MIN5404. Furthermore, the royalty agreement extends north and south of MIN5404 where previously existing tenement EL3211 (New Holland Mining) overlaps with EL3539 (Perseverance). This is shown in Figure 2.

There are no Native Title issues relevant to MIN5404.

The environmental bond is currently set at A$5,131,523 and is reviewed annually with the Department of Primary Industries in Victoria. Rehabilitation is undertaken progressively at the Fosterville Gold Mine but the environmental bond is only reduced on establishment of the rehabilitation which is not considered to have occurred until at least 5 years after rehabilitation has occurred.

The Fosterville mine is located near areas of moderate environmental significance (Mt Sugarloaf Native Conservation Reserve), established productive farmland and is adjacent to the locally significant Campaspe River.

The Fosterville Gold Mine is operating under a Work Plan approved in April 2004 under section 44(1) of the Mineral Resources Development Act. The approval, concerning MIN5404 (formely ML1868), MIN4456 and MIN4887, was given by the Minister of Environment and Water at that time to Perseverance Exploration Pty Ltd. Work Plan Variations are submitted where significant changes from the Work Plan exist. The latest WPV was approved in November 2007.

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

The Fosterville area is flat to very gently undulating with a range of low, rolling hills about 2 kilometres to the west and the Campaspe River about 2 kilometers to the east. On ML5404 natural surface elevations range from 150 to 185 metres above sea level (5150RL to 5185RL mine grid). Vegetation in the area ranges from native forest to established grazing pasture.

The Fosterville Gold Mine has ready access via two separate sealed roads and a variety of all weather un-sealed roads linking to regional highways. The regional centre of Bendigo is approximately 20 kilometers away and has a population of 95,000 people which provides a source of skilled labour. The area has a mediterranean climate with hot, dry summers and cool winters. The climate statistics from the nearby Bendigo Airport weather station for the period 1991-2007 are listed in Table 4 below.

Statistics	Jan	Feb	Mar	Apr	May	Jun	Jul	Aug	Sep	Oct	Nov	Dec	Annual
Temperature
Mean maximum
temperature (°C)	29.2	29.4	25.7	21.2	16.6	13.2	12.5	14.3	16.6	20.1	24	27	20.8
Mean minimum
temperature (°C)	13.8	14.1	11.3	7.5	5.2	3.4	2.4	2.5	4.4	6.2	9.2	11.4	7.6
Rainfall
Mean rainfall (mm)	32.2	27.8	21.6	27.1	49.6	51.7	49.6	45	50.8	43.6	43.4	40.9	482.3
Decile 5 (median)
rainfall (mm)	23.4	22.2	11.2	23.5	39.9	39.8	49.2	41	40.4	45.5	34.2	30.4	459.7
Mean number of days
of rain = 1 mm	4.1	3.1	3.2	4.1	6.5	7.9	8.4	7.4	8	5.8	5.5	4.6	68.6

Table 4: Climate statistics from the Bendigo Airport weather station, 1991-2007 (Bureau of Meteorology, 2008).

Power is supplied to the site via a terminal station that was constructed by Perseverance in 2005. This station is connected to the 220kV transmission line that runs from Bendigo to Shepparton which traverses the southern end of ML5404 approximately 1.5 kilometres south of the processing plant. There is a connection agreement in place with SP Ausnet who manages the transmission and distribution network. To improve the security of water supply, an agreement was reached for the supply of waste water from the Bendigo sewerage treatment facility. A pipeline was commissioned in April 2005 which has the capacity to supply approximately 2000 ML annually which comfortably exceeds the current plant useage of approximately 1000 ML per annum. The agreement was for an initial 10 year term with 2 options of a further 10 years each on written request.

All other site infrastructure is in place and approved in the Work Plan established in April 2004.

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6. History

Gold was first discovered in the Fosterville area in 1894 with activity continuing until 1903 for a total of 28,000 ounces of production. Mining in this era was confined to near-surface oxide material.

Aside from a minor tailings retreatment in the 1930’s, the field lay dormant until the 1988 when Bendigo Gold Associates again recommenced gold production at Fosterville from the reprocessing of tailings. By 1989 this program had come to an end and exploration for oxide resources commenced. The leases were then acquired by Brunswick who continued exploration and in 1991 started heap leaching ore derived from shallow oxide open pits. After six months of production, Brunswick went into receivership as a result of the failure of another operation. Perseverance bought the operation from the receivers and continued the oxide heap leach operations. Perseverance continued to produce between 25,000 oz to 35,000 oz per annum until the cessation of the oxide mining in 2001. Between 1988 and 2001, a total of 240,000 ounces of gold were poured (Roberts et al, 2003).

In 2001, Perseverance underwent a significant recapitalisation and the focus of the company changed to developing the sulphide resource. A feasibility study investigating a combined open pit and underground mining operation feeding 0.8Mtpa of sulphide ore to a BIOX® processing plant was completed in 2003. Work on the plant and open pit mining commenced in early 2004. Commercial sulphide hosted gold production commenced in April 2005 and up to the end of December 2007 had produced 263,900 ounces of gold. Underground development commenced in March 2006 with first production recorded in September 2006. Table 5 below gives a breakdown of open cut and underground mined tonnes and grade since the commencement of the sulphide operation.

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		2004	2005	2006	2007
Open Cut Mined	Tonnes (t) Grade (g/t)	52,000 3.6	517,000 5.6	1,084,000 3.4	423,000 2.3
Underground Mined	Tonnes (t) Grade (g/t)			36,000 4.8	376,000 4.2
Total Mined	Tonnes (t) Grade (g/t)	52,000 3.6	517,000 5.6	1,120,000 3.4	799,000 3.2

Table 5: Mined production data for Fosterville for the period 2004-2007.

On 29th October 2007, Perseverance announced that it had entered into an agreement with Northgate Minerals Corporation (Northgate) to acquire the company via a Scheme of Arrangement. This agreement was ratified by Perseverance’s shareholders and optionholders on the 18th January 2008 with full control passing to Northgate in February 2008.

7.Geological Setting

7.1 Regional Geology

The Fosterville Goldfield is located in the Bendigo Zone in the Lachlan Fold Belt. The host rock lithologies in this zone are dominated by a sequence of folded and faulted Ordovician turbidites which were subsequently deformed in the Late Ordovician (450-430 Ma) Benambran Orogeny. The sediment pile was deformed under east-west compression resulting in the formation of north-south folds. As this process continued and fold limbs steepened, a series of west-dipping reverse faults progressively developed. This generation of faults is interpreted to have a listric geometry and were likely conduits for ascending mineralised fluids (Figure 3).

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Figure 3: Schematic cross-section (EW) through the upper crust in the western Lachlan Fold Belt.

There are two broad periods of gold mineralisation indicated by radiogenic dating across the western Lachlan Fold Belt. The first of these is associated with quartz vein hosted mineralisation and is concentrated from 425 - 400 Ma. This period and style of mineralisation extended from Stawell in the west to Bendigo - Wattle Gully in the east. Fluid inclusion studies indicate that this style of mineralisation formed at temperatures between 320°C and 400°C and at depths of 5-10 kilometres (Roberts et al, 2003).

The Bendigo Zone was intruded by two granitic suites during the early Devonian and again in the late Devonian. These events appear linked to the second phase of mineralisation which occurred between 380 - 365 Ma and extended from Ballarat in the west to the Woods Point - Walhalla belt in the east. The Fosterville mineralisation appears to have formed during this phase. Mineralisation from this second phase can manifest in a range of styles from quartz-carbonate vein hosted free gold through to sulphide hosted refractory gold in association with arsenopyrite, pyrite and stibnite (Roberts et al, 2003).

Deep weathering and erosion in the late Tertiary resulted in the development of a regional laterite profile with weathering locally to 50 metres depth.

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7.2 Deposit Geology

The Fosterville Goldfield is hosted by a turbiditic sequence of interbedded sandstones, siltstones and shales interpreted as having formed through a regime of meandering submarine channels. The sequence is dominated by shale topped sands ranging from 0.2 to 1.5 metres thick, with lesser amounts of massive sandstone, shale and black shale (Roberts et al, 2003).

The sequence was metamorphosed to sub-greenschist facies and folded into a set of upright, open to closed folds. This folding resulted in the formation of an axial planar and radial cleavage, best developed in fold hinges. It also resulted in the development of bedding parallel laminated quartz veins, preferentially forming in shales at or close to the contact with sandstone units. These laminated quartz veins are visually similar to those that carry free gold at Bendigo (20 kilometres away), however at Fosterville they are effectively unmineralised.

Mineralisation at Fosterville is controlled by a series of late brittle faults which are often bedding parallel and follow pre-existing laminated quartz vein, however they do also crosscut bedding to link bedding parallel faults across folds. They are generally steeply west dipping reverse faults with a series of moderately dipping reverse splay faults formed in the footwall of the main fault. The splay faults are generally confined to the zone within 100 metres east of the main fault. In the current mining areas in the Central Area, the main fault is the Fosterville Fault with the Phoenix Fault being the most important splay fault in terms of identified mineralisation (Figure 4).

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Figure 4: Cross-section though Phoenix and Ellesmere Shoots showing mineralization domains, faults, bedding formlines and the location of hanging wall black shale.

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Fluid inclusion studies at Fosterville indicate that mineralisation formed at temperature ranges of 140°C to 385°C and at depths of 2.6 to 5.7 kilometres.

At Fosterville, several felsic dykes up to 10 m wide have intruded the Robbin’s Hill area. A series of thin, deuterically altered basaltic dykes post-date all significant faulting and mineralisation, and one has been dated at 170 ± 5 Ma by Ar³⁹/Ar⁴⁰ (Vandenberg et al, 2000).

Neogene erosion followed by valley infill and weathering has resulted in local clay - conglomerate alluvial channels and complete oxidation to about 5120mRL (30 to 40 metres below surface). Immediately below the base of complete oxidation is a 10 to 15 metre thick zone of partial oxidation of sulphide minerals. Feldspar destruction and partial carbonate dissolution extends from the base of oxidation to about 5000mRL.

8. Deposit Types

Gold mineralisation at Fosterville is relatively homogenous with only one deposit type present. There are minor variations in the host rock type and structural setting. Fosterville-type deposits form a sub-group of orogenic gold deposits that are typified by gold occurring in fine grained arsenopyrite and / or pyrite disseminated in country rocks as a selvage to faults or veins. Fosterville-type deposits and classic vein-hosted deposits are effectively end members with many orogenic gold deposits displaying features of both.

Primary mineralisation at Fosterville is controlled by late brittle faulting. These late brittle faults are generally steeply west dipping reverse faults with a series of moderately west dipping reverse splay faults formed in the footwall of the main fault. Primary gold mineralisation occurs as disseminated arsenopyrite and pyrite forming as a selvage to veins in a quartz – carbonate veinlet stockwork. The mineralisation is structurally controlled with high grade zones localised by the geometric relationship between bedding and faulting. Mineralised shoots are typically 4 to 15 metres thick, 50 to 150 metres up / down dip and 300 to 1500 metres+ down plunge. These sulphide bodies are the primary target for exploration activities, especially where there is potential for grades in excess of 3 g/t Au (ie. above likely underground cut-off grades).

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Within the oxide zone, there has typically been minor re-mobilisation of gold into the immediately surrounding country rocks which has resulted in a small increase (about 50%) in the width of mineralisation and consequent reduction in gold grade. There is no evidence of a wide spread high grade supergene zone immediately below the water table. There is no current focus on exploring for additional oxide resources.

9. Mineralisation

Primary gold mineralisation at Fosterville occurs as disseminated arsenopyrite and pyrite forming as a selvage to veins in a quartz – carbonate veinlet stockwork which is in turn controlled by the late brittle faults.

The arsenopyrite occurs as fine grained (0.05 to 1 mm), acicular needles with no preferred orientation. The disseminated pyrite associated with gold mineralisation occurs as crystalline pyritohedrons 0.1 to 2 mm in size. Electron microprobe analyses and metallurgical testwork indicates that the arsenopyrite contains 100 ppm Au to 1000 ppm Au and the auriferous pyrite 10 ppm Au to 100 ppm Au. Approximately 80% of the gold occurs in arsenopyrite, with the remaining 20% hosted by pyrite. Silver grades are low at Fosterville, usually about one tenth of the gold grade.

Framboidal aggregates and laminations of pyrite up to 20mm are common, especially in black shale units. The framboidal pyrite is diagenetic and is not auriferous.

Antimony mineralisation occurs as very coarse grained overgrowths of stibnite up to 20 centimetres across replacing late quartz - carbonate veins. The stibnite appears to contain 1 ppm Au to 10 ppm Au, however there is usually a high grade (15 g/t Au to 40 g/t Au) arsenopyrite and pyrite mineralisation occurring as a selvage to the quartz - stibnite veins. Antimony mineralisation appears to be restricted to splay faults.

The quartz – carbonate veinlet stockwork comprises a network of tension gash type quartz – carbonate veinlets which have formed perpendicular to the walls of the brittle faults and quartz – carbonate veinlets formed on minor slip planes parallel to the brittle faults. Further movement on the minor slip planes offsets the tension gash veinlets giving rise to a range of geometries from planar through to highly erratic. The quartz – carbonate veinlets are barren but have selvages of disseminated, fine grained arsenopyrite – pyrite. Where the stockwork is well developed, mineralisation selvages merge forming a solid body of mineralisation. On the margins of the stockwork the mineralisation occurs as a discrete selvage about 10 times the width of the veinlet on which it is centred.

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9.1 Controls on Primary Mineralisation

Primary mineralisation is structurally controlled with high grade zones localised by the geometric relationship between bedding and faulting. Mineralised shoots are typically 4 to 15 metres thick, 50 to 150 metres up / down dip and 300 to 1500 metres+ down plunge. Gold grades are relatively smoothly distributed with both extremely high values and extremely low values rare.

9.1.1 Fosterville Fault Zone

9.1.1.1 Central and Hunts Areas

A culmination in the plunge of the footwall syncline occurs at about 8800mN and this culmination is mirrored in the plunge of the mineralisation shoots. To the north of 8800mN, the footwall syncline and mineralisation shoots plunge gently to the north. Similarly south of 8800mN, the footwall syncline and mineralisation shoots plunge gently to the south.

There are three general fault – bedding geometric relationships (Figure 5).

Immediately adjacent to the main hangingwall fault, the splay faults traverse through east dipping beds which are oblique on both the hangingwall and the footwall of the splay fault. This part of the splay fault, referred to as an oblique – oblique segment, is usually moderately mineralised. Oblique – oblique segments create fracture porosity in both the footwall and the hangingwall of the splay fault.

Moving up dip beyond the intersection of the axial plane of the syncline in the footwall and the splay fault, there is a zone with oblique east dipping beds in the hangingwall of the splay fault and parallel west dipping beds in the footwall. Oblique - parallel segments create fracture porosity only in the hangingwall of the splay fault and typically host the best developed mineralisation.

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Further up dip the splay fault passes under the axial of the syncline in the hangingwall so that there are concordant west dipping beds in both the hangingwall and footwall of the splay fault. Parallel – parallel segments are usually weakly and discontinuously mineralised because the bedding parallel splay fault doesn’t create fracture porosity for the transmission of mineralising hydrothermal fluids.

At the point where the splay fault passes under the axial of the syncline in the hangingwall, the mineralising hydrothermal fluids appear to have used fracturing in the hinge of the syncline to traverse up to the overlying fault which may be either the hangingwall fault or the next splay up sequence.

The geometry of the brittle faults also provides some secondary control on localising mineralisation with shallower dipping sections and left stepping segments providing favourable dilational zones.

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Figure 5: Schematic cross section showing the relationship between mineralisation, faulting and bedding orientation.

9.1.1.2 Wirrawilla – Daley’s Hill Area

South of 7000mN, the Fosterville Fault and the Western Bounding Fault (WBF) merge and an anticline is observed in the footwall so that the Fosterville Fault / WBF is in a parallel – parallel position. Consequently, there is no mineralisation on the Fosterville Fault / WBF.

Alternatively, mineralisation in this area occurs as shoots on individual fault strands within the Harrier Fault complex, a complex of spay faults that is exposed in the Harrier pit and has been traced south to Daley’s Hill. The same relationships between faulting and parallel or oblique beds seen in the Central area splay faults are also seen within the Harrier Fault complex. The Harrier Fault Complex plunges gently to the south and is open south of 4700mN. Gold mineralisation in the Harrier complex tends to be narrower, more complex and less continuous than in the Central Area.

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Very rare primary free gold has been discovered on a single fault approximately 80 metres east of, and striking parallel to, the Daley’s Hill Fault. This primary free gold mineralisation overprints earlier sulphide mineralisation, is high grade (10 – 30 g/t Au), narrow (1.5 m to 3.0 m true width) and is limited by drilling to a strike extent of 80 metres and is open down dip. Sulphide hosted gold is still dominant in this zone. The significance of this style of mineralisation is not yet known.

9.1.2 Robbins Hill Area

Although the Robbins Hill area is less intensely explored at depths below 100 metres, fault controlled primary gold mineralisation exhibits structural and lithological controls similar to the Fosterville Fault Zone. Faults are clearly the primary control on mineralisation with lithology providing secondary control. However, it is apparent that the controls recognised in the Central Area are not directly applicable as mineralisation is observed in parallel – parallel positions. The mineralised faults generally strike within 20° of north and dip moderately to steeply to the west.

Dyke associated gold mineralisation occurs, mainly within 2 metres of the dyke contacts. The dykes are generally sub-vertical and have often intruded faulted out fold hinges. Higher grade zones are controlled by the intersection of fault controlled mineralisation with the dykes.

9.2 Controls on Oxide Mineralisation

Minor re-mobilisation of gold into the immediately surrounding country rocks has resulted in a small increase (about 50%) in the width of mineralisation and consequent reduction in gold grade. There is no evidence of a wide spread high grade supergene zone immediately below the water table.

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Other elements have been more significantly affected by weathering processes. Dissolution of sulphur by oxidising groundwater above the water table has effectively removed all sulphur from the oxide zone. Arsenic has been strongly remobilised over a zone five to ten times the width of mineralisation. The greater width of anomalous arsenic values in the oxide zones makes arsenic soil geochemistry a very useful tool for finding exposed gold mineralisation.

10. Exploration

10.1 Pre-1992

Modern exploration commenced at Fosterville during the 1970s. Apollo drilled three HQ diamond holes in what is now the Hunts area. Noranda drilled three HQ diamond holes in the Daley’s Hill area. None of these holes have been included in the drilling database in the due to uncertainty in their collar locations.

From 1987 to 1991 Bendigo Gold Associates and later Brunswick drilled 488 RC holes and 6 HQ diamond holes targeting oxide mineralisation on the Fosterville Fault and the Robbins Hill area. This program resulted in the development of a heap leach operation which commenced in 1991.

Brunswick also completed a 100 metres by 20 metres soil geochemistry grid across the project area and as far west as the Sugarloaf Range. The soil geochemistry was very effective at defining gold mineralisation except where alluvial cover exceeded about two metres. Two preliminary IP lines were completed with mixed results.

10.2 1992-2001

On acquiring the Fosterville mining lease in 1992, Perseverance (through a drilling contractor) started RC drilling for further oxide resources and reserves using a combination of cross over and face sampling hammers.

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In late 1994, while continuing to explore for oxide mineralisation, Perseverance began to drill for sulphide mineralisation on the Fosterville Fault potentially amenable to open cut mining. This drilling was almost entirely RC using a face sampling hammer with minor diamond drilling for metallurgical and geotechnical purposes and extended from 6000mN to 10700mN. Most of the drilling was completed by 1997 with minor infill drilling continuing to 1999.

Section spacing was either 25 or 20 metres except in two small zones in the Falcon and Ellesmere areas where 12.5 metres sections were drilled. Vertical spacing was nominally 20 metres. This drilling programme was generally restricted to within 100 metres of surface, extending to a vertical depth of 150 metres below surface in the Central North area, reflecting the perceived limits of open cut mining. The data from this drilling program formed the basis of the 1997 Sulphide Project Feasibility Study which was later updated in 2000 (Perseverance, 1997; 2000).

Two deep diamond holes, SPD7 and SPD8 were also drilled. SPD7 was drilled beneath the Central Ellesmere pit and intersected 53.8 metres @ 1.97 g/t Au (hole abandoned in mineralisation) from 382.0 metres, while SPD8 was drilled to 450.0 m below Central North intersecting only 2.0 metres @ 0.58 g/t Au on a splay fault some 60 metres to the east of the Fosterville Fault.

A 25 metre by 25 metres gradient array IP survey was conducted in the Robbins Hill area in 1997. This survey did not conclusively define gold mineralisation, however it was successful in mapping carbonaceous shales and alluvial channels.

10.3 2001-Present

The current drilling program on the Fosterville Fault Zone commenced in July 2001 and is ongoing. For the majority of this period, the surface drilling activities have been conducted by Silver City Drilling Pty Ltd (drilling contractor) and the underground drilling activities have been conducted by Deepcore Pty Ltd (drilling contractor). Resource definition holes are usually drilled with RC pre-collars and NQ2 diamond tails. The sectional spacing ranges from 200 meters to 50 metres with the vertical spacing of intersections usually 50 metres.

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A small number of RC only holes have been drilled where the target was shallow and exploratory. Once definitive targets were defined by this type of drilling, the drilling methods changed to the those used for resource definition drilling described above.

The change in drilling methods to largely oriented diamond core, intensive re-mapping of old oxide pits and a change in logging methods to collect detailed grain size data allowing sequence stratigraphic analysis allowed much more detailed and robust geological models. These geological models allowed a better understanding of the controls on gold mineralisation which in turn resulted in the better targeting and more efficient use of drilling.

The post-2001 exploration has resulted in the discovery and definition of the Phoenix, Wirrawilla and Farley’s deep zones. In addition the Falcon, Ellesmere and Harrier zones have been extensively extended. Modest additions to resources have been made at the Daley’s Hill, Sharkey’s and Hunts deposits.

Two further IP surveys were completed in 2001 and 2005. The 2001 survey consisted of four lines of 50 metre nodes over the central area. This survey was designed to define gold mineralisation at depths of between 50 to 250 metres. The data was inverted to make a model in real space. Anomalies were defined along the Fosterville Fault zone, but the 50 metre node spacing meant that the survey resolution was unable to distinguish the carbonaceous shale in the hangingwall of the Fosterville Fault from mineralisation in the footwall of the Fosterville Fault. In 2005 another four IP lines were completed across the northern end of the Fosterville Goldfield, covering the Sugarloaf geochemical anomaly, the Fosterville Fault Zone and the Robbin’s Hill area. This survey defined weak anomalies over the Sugarloaf geochemical anomaly and the strike projection of the Fosterville Fault Zone north of MIN5404.

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11. Drilling

11.1 Pre-1992

Section 10.1 described the work undertaken from the commencement of modern exploration in the 1970’s through to 1992. For the period 1987-1991 (Bendigo Gold Associates, then Brunswick), drilling used a cross over type hammer and significant downhole contamination is suspected for some holes below the water table where the rig was unable to produce dry samples. Although this drilling method is likely to have produced a relatively poor sample quality, the drilling was confined to oxide pits that have subsequently been mined and so is not relevant to the resources and reserves included in this report.

11.2 1992-2001

On acquiring the Fosterville mining lease in 1992, Perseverance (through a drilling contractor) started RC drilling for further oxide resources and reserves using a combination of cross over and face sampling hammers. These holes used the CN, CEL, CEN, DH and HAR prefixes.

In late 1994, while continuing to explore for oxide mineralisation, Perseverance began to drill for sulphide mineralisation on the Fosterville Fault potentially amenable to open cut mining. The 1997 Feasibility Study drilling was almost entirely RC with minor diamond drilling for metallurgical and geotechnical purposes and extended from 6000mN to 10700mN. Most of the drilling was completed by 1997 with minor infill drilling continuing to 1999. Holes from this programme have the SP (sulphide project), CN, CEL(D), CEN(D), GT or HAR(D) prefixes, the ‘D’ denoting holes with a diamond tail.

All the RC holes used face sample hammers. After 1996, if the sample was unable to be kept dry the hole was finished with an NQ diamond tail. The RC samples were initially sampled using a PVC ‘spear,’ however after 1995 all RC holes were sampled using a 12.5% riffle splitter. All holes were completely sampled, the RC holes as two metre composites and the diamond tails as one metre intervals of split core.

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As part of the 1997 Feasibility Study several FO holes with long, high grade intersections were twinned with RC holes drilled with a much bigger compressor and a face sample hammer resulting in dry samples. These twin holes demonstrated that there was significant downhole contamination in the FO holes (Perseverance, 1997). As a result, the FO holes were only used for estimating oxide resources and reserves where it is assumed that dry samples were recovered and downhole contamination was not an issue.

Open hole downhole surveys were completed on all holes at 30 metre intervals except for a small number of holes which collapsed before a survey instrument could be lowered down the hole. The vast majority of holes were drilled from the west towards the east, generally intersecting mineralisation at 50° to 80°. Most sections include at least one hole drilled towards the west as a check on the geological interpretation.

The Fosterville Mine Surveyor used a Total Station Instrument to run a complete digital survey of the topography for any areas where drilling and later resource evaluation was planned to take place. Spot heights were measured at suitable intervals where easting, northing and RL are noted. Closer spaced measurements were taken around noticeable highs and lows in the topography. These spot heights were then triangulated using Minsurv software to construct a Digital Terrain Model (DTM). This DTM was used in all resource/reserve estimates at Fosterville. The spot heights were measured to an accuracy of ± 1 cm at spacings of approximately 20 metres.

11.3 Current Drilling Methods

The current drilling programme on the Fosterville Fault Zone commenced in July 2001 and is ongoing. For the majority of this period, the surface drilling activities have been conducted by Silver City Drilling Pty Ltd (drilling contractor) and the underground drilling activities have been conducted by Deepcore Pty Ltd (drilling contractor). The SPD holes were drilled with RC pre-collars and NQ2 diamond tails. The diamond tails commenced at least 20 metres before the hanging wall fault so that all mineralisation was intersected by the diamond tail.

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The RC pre-collars were generally 150 metres to 200 metres deep and the diamond drilling was double tube wireline drilling. In addition, 15 wedges have been drilled from 10 parent holes.

Collar locations are surveyed using the same technique as prior to 2001 (see section 11.2 above).

The direction of the RC pre-collars was controlled to some degree by the use of a stabiliser rod, the relative size of the bit compared to the rods and by the weight on the hammer. In general, holes shallower than 70° tended to lift, whereas holes steeper than 75° tended to drop. With experience deviation in the pre-collar was restricted to less than 1° in 100 metres. Directional navi drilling was occasionally used to keep holes on target where the RC pre-collar deviated significantly. Down hole surveys were carried out using a single shot Eastman camera at 25 metre intervals in the pre-collars (every 50 metres inside the rods as the hole was drilled and the intervening 25 metre intervals open hole after the pre-collar was completed) and at 30 metre intervals in the diamond tails. As a check on the validity of the single shot surveys six holes were surveyed at 6m intervals using an EMS (electronic multi-shot) tool (see Section 14.3 below).

The drillhole traces are currently calculated using the ‘semi tangent’ de-surveying algorithm in Minesight. This method is suitable for deeper RC holes, which have more than 2 downhole surveys. The ‘fit-spine’ algorithm was previously used because it dealt well with RC holes that have only one or two surveys near the top of the hole and also because this algorithm was used historically at Fosterville.

The NQ2 diamond core has generally been drilled using 6 metre core barrels. The core was oriented using a spear about every 30 metres. Approximately 50% of the core was able to be directly oriented from the spear marks, the orientation of about another 45% of the core was inferred using a reference plane (either bedding or cleavage) whilst the remaining 5% or so was unable to be oriented. Two new core orientation devices were trialled on about 30 holes during the last quarter of 2002. The systems trialled were the ezimark and ballmark systems, both of which should give an orientation every run, although due to broken ground this is usually more like every third run. The ballmark system was chosen due to better repeatability between runs and has been used routinely since the beginning of 2003.

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All of the RC pre-collars were sampled as four metre composites of 6.25% splits from a riffle splitter. However, mineralised portions of drill holes were later riffle split to one metre lengths to better define gold mineralisation. Sampling of RC pre-collar holes after May 2004 was changed such that two metre composite samples were exclusively conducted throughout all drill programs.

Sieved chips from the RC pre-collars were logged in two metre intervals for lithology, weathering, alteration, % quartz, colour and recovery. The information was entered directly in the field into a hand held computer (IPAQ) and downloaded to the database. The downloading procedure has in built checks to prevent interval overlap, range checking etc. After downloading the entire log is printed for hand plotting and as a hard copy record.

The diamond core is roughly oriented at the rig by the drillers before transport to the core shed. At the core shed the core is washed, oriented, digitally photographed, recovery and RQD measured, geologically logged, marked for sampling, sampled and samples dispatched. Geotechnical logging occurs on an as needs basis. The remaining core is stored in the core farm behind the core shed. The geological logging involves directly entering observations on sediment grainsize, lithology, planar and linear structural observations (as alpha, beta and gamma), mineralisation, alteration and quartz veining into a hand held computer. In addition, sample numbers are also recorded into the hand held computer. When logging is complete the hand held computer is downloaded into the database with the usual automated error checking, a list of samples printed as a cutting sheet and the entire log printed for hand plotting and as a hard copy record.

In the diamond core all visible sulphide mineralisation, quartz vein stockworks and laminated quartz veins plus at least two metres of apparent waste was sampled. Samples were cut to geological boundaries and ranged in length from 0.3 m to 2.0 m, with a preferred length of 1.0 m. The core was halved along the plane of orientation using a diamond saw and the upper half of the core dispatched for analysis and the lower half returned to the core tray in its original orientation. The PQ core was sampled by cutting a sliver equivalent in volume to ½ NQ2 core from the top of the core.

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The strategy for underground diamond drilling of the Falcon and Phoenix ore zones is to infill the exploration drilling intercepts to a notional 25 metre x 25 metre grid spacing (or tighter if required) prior to the mining of underground development. Underground diamond drill core samples used in the Falcon and Phoenix resource estimations were either NQ2 or LTK60 in size, depending on the drill rig used.

Figure 6: Underground diamond drilling traces (green) in footwall of Phoenix and Falcon ore zones.

The nominal progression of drilling is from initial surface exploration drilling, through 100 metre by 100 metre and then 50 metre by 50 metre. Near surface mineralisation is then further infilled to 25 metre by 25 metre to allow pit design. Open pit grade control drilling is inclined RC holes drilled 5 metres apart on either 10 or 12.5 metre sections to a maximum depth of 30 metres. The open pit drilling, sampling and logging methods are the same as exploration RC drilling. Underground mineralisation is infilled to 25 metres by 25 metres or tighter if required by underground diamond holes. Strike drives on 20 metre levels are sampled every round (four metres).

Based on drilling results, interpretations are made in cross-section and level plan to interpret the true thickness of the mineralised zones with geological solids subsequently generated from these interpretations. Further detailed discussion on this process is contained in Section 17 under each of the modelled areas.

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Sulphide Series		Holes	Comments
BGL001	BGL104	21	1990-2004 RC Hydrological holes
CEL001	CEL124	96	1997- Open pit sulphide holes
CELD (3)	CEND (4)	7	1998 PQ core for JK drop tests
CELD020	CELD103	23	1997-8 Diamond Tails from RC wet drilling
CELD051	CELD058	8	1996 Metallurgical Diamond holes
CEM100	CEM105	6	1994 RC metallurgical holes
CEN001	CEN124	80	1997- open pit sulphide holes
CEND019	CEND103	23	1997-8 Diamond Tails from RC wet drilling
CEND038	CEND065	9	1996 Metallurgical Diamond holes
CN100	CN248	149	1994 RC exploration holes
CNM1		1	1995 Metallurgical RC hole
ELRC0001	ELRC0949	912	2005-2007 RC holes 7500mN – 8425mN
FARC0001	FARC0825	825	2005 RC holes 8615N – 8800N
FDD-14	FDD-40	6	1990 Brunswick Holes
FOS056	FOS214	3	1998-2000 RC exploration holes
GT1	GT4	8	1996 HQ diamond Geotechnical holes
HAR1	HAR65	61	1997-99 RC exploration holes, Harrington’s Hill area
HARC0001	HARC0235	218	2006 – 2007 RC holes 6900mN – 7315mN
HARD1		1	1996 PQ diamond metallurgical hole
SP001	SP372	299	12/1994-10/1996 RC Holes to 5100 RL
SPD010	SPD470	354	2001-05 Definition and delineation of Falcon, Phoenix, Ellesmere and Harrier.
SPD-1	SPD-9	9	1995 Diamond exploration holes
ST009	ST172	50	Sterilisation holes, do not intersect mineralisation
UD001	UD190	190	2006 – 2007 Underground diamond holes
	Total	3,358	Holes

Table 6: Drill hole prefixes for all drilling on the Fosterville Fault Zone between 6000mN and 10000mN.

12. Sampling Method and Approach

Prior to Perseverance acquiring the project in 1992, RC drill samples were collected at one metre intervals using ‘spear’ sampling. Although this sampling method is likely to have produced a relatively poor sample quality, the drilling was confined to oxide pits that have subsequently been mined and so is not relevant to the resources and reserves included in this report.

From the acquisition of the project by Perseverance in 1992 through to the present, all RC drilling through mineralisation has been collected at one metre intervals and sampled as two metre composite samples. Prior to 1995, samples were collected using ‘spear’ sampling. Since 1995 all RC holes have been sampled using a riffle splitter split to either 12.5% or 6.25% depending on the hole diameter. After 1996, if the sample was unable to be kept dry the hole was finished with an NQ diamond tail. In the central area, spear samples comprise 16% of all mineralised samples and 28% of all mineralised RC samples. All RC holes were completely sampled.

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In the diamond drill core, all visible sulphide mineralisation, quartz vein stockworks and laminated quartz veins plus at least two metres of apparent waste either side was sampled. Samples were cut to geological boundaries and within a length range of 0.3 metres to 2.0 metres, with a preferred length of 1.0 metres. The core was halved along the plane of orientation using a diamond saw and the upper half of the core dispatched for analysis and the lower half returned to the core tray in it’s original orientation. The PQ core was sampled by cutting a sliver equivalent in volume to ½ NQ2 core from the top of the core. Recovery of diamond drill core is acceptable with >98% recorded for the drillholes incorporated into the Central Area resource models.

In underground sampling, an attempt is made to sample every round (4 metre nominal advance) in the ore drives where safe to do so. Sample intervals are chosen based on lithology, alteration and mineralisation, and are a minimum of 0.3 m and a maximum of 1.5 m in length. Mapping data that was collected at the same time as the samples is used to validate the sample results.

A program of duplicate sampling was undertaken on the Phoenix 5020 level in 2007, where a field duplicate was taken for every sample collected. The results for 174 pairs of samples show reasonable repeatability with an R² of 0.803. This study covered the underground face sampling method used throughout the mine since ore driving commenced in late 2006 and the area represented typical geology in terms of mineralisation and geometry. On this basis, it is reasonable to apply this level of confidence to all of the face sampling included in the Central Area resource models.

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Figure 7: Face sample field duplicate results

All remaining diamond drill core is stored on site in racks within the fenced and gated core handling facility. Assay sample pulps are also returned from the laboratory and stored at the core handling facility.

RC samples from previous grade control programs were kept at an onsite depot for approximately 3 months after the receipt of final assay results. This allowed time for any re-sampling that may be necessary. The plastic sample bags photo-degrade rendering re-sampling impossible after 6 to 12 months and presenting an environmental hazard from windblown plastic, therefore the sample bags are disposed of as part of routine site rehabilitation works. Exploration RC pre-collar samples were collected in hessian sample bags since 2005 and similarly retained for a 3 month period at the drill sites. Hessian was chosen as it poses less of an environmental hazard and allows for mechanical rehabilitation of drill sites.

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13. Sample Preparation, Analyses and Security

13.1 Elements

Elements analysed are shown in Table 7.

Element	Selection of samples
Au	All fresh samples by 40g Fire Assay to Dec 2004 and then by 25g Fire assay, all oxide samples by 25g aqua regia digest
As	All, since August 1995
S	For all Au values over 0.5g/t August 1995 to May 2001, then all samples
Sb	For all Au values over 0.5g/t August 1995 to May 2001, then all samples
NCC (non-carbonate carbon)	For all Au values over 0.5g/t August 1995 to May 2001 From 2006, all sample intervals that are in dark shales or carbon bearing zones.
TGC (total graphitic carbon)	Selected samples between June 2002 and August 2003
TOEC (Total Organic and Elemental Carbon) – Equivalent to NCC	Selected RC pre-collar samples and most diamond samples since 1996
CO3	selected samples only
Cu, Pb, Zn, Ag, Fe, Mn, Mg, Bi, Ca, Cd, Ce, Co, Cr, K, Mo, Na, Nb, Ni, P, Sr, Ti, V, Y Ag, As, Bi, Ca, Cu, Fe, K, S, Sb	All between May 2001 and Feb 2006, by ICP-AES, 5 gm HF mixed acid digest ICP3E, AMDEL All since Feb 2006 by ICP-AES, modified triple acid digest, OSL

Table 7: Analysed elements by method and time period.

The elements important to the metallurgical oxidation process are Au, S and Sb in decreasing importance. Arsenic is modelled for environmental reasons. NCC (non-carbonate carbon) is of importance in the CIL stage as graphite is preg-robbing of gold in solution.

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13.2 Description of Analytical Techniques

All of the gold analyses used in the sulphide resource model in the 2000 Sulphide Feasibility Study were fire assays of a 40g charge carried out by ALS at Bendigo, a commercial laboratory (non-accredited). The other elements were analysed by a variety of techniques at a variety of laboratories. A full program of repeats, standards and inter-laboratory check sampling was conducted on the gold analyses.

For the 2001 – 2004 NQ2 SPD diamond drilling campaign, gold analyses were determined by fire assay of a 40g charge by AMDEL in Adelaide, a commercial laboratory (ISO 9001 accredited). A 30 element suite including As, S and Sb was analysed by ICP-AES from a separate 5g charge following HNO3 / HF digestion. From November 2002 to August 2003 TGC (total graphitic carbon) was analysed on a selective basis. A full program of repeats, standards and inter-laboratory check sampling was conducted on the gold analyses.

Since 2005, On Site Laboratory Services (OSL), a commercial laboratory based in Bendigo, was the primary provider of analytical services to the project. The OSL Bendigo laboratory is currently not accredited, however it is working towards ISO 9001 accreditation which is anticipated to be in place by April 2008. Following sample drying, OSL use a combined crusher and mill to pulverise the entire sample to a nominal 95% passing 75µm. A 25 g sub-sample is analysed for gold by fire assay with an AAS finish. A 0.5 g sub-sample of the pulp is digested in a HNO3 / HCl digest and then analysed for Ag, As, Bi, Ca, Cu, Fe, K, Sb and S by ICP-AES. A full program of repeats, standards and inter-laboratory check sampling was conducted on the gold analyses.

An audit of the OSL facility was completed for Perseverance by an external consultant during 2007 (Stewart, 2007). This audit found that OSL’s procedures were adequate and presented no major risk to the resource estimate. There were areas for improvement identified with the following corrective actions taken during the second half of 2007:

1.

Temperature variation within the drying oven is now being measured and recorded.

2.

Sizing analysis for all pulps is now being conducted and recorded.

3.

Calibration of scales is now being recorded and documented.

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Work undertaken by employees of the company was limited to core logging and the markup, cutting and bagging of samples. All other sample preparation and analysis was conducted off-site at the commercial laboratories.

13.3 Sample and Data Security

13.3.1 Sample Security

Sample security information has not been recorded over the history of the project. However, to the best of the authors’ knowledge, the methods of sample storage and transport have remained largely unchanged throughout the life of the project.

Samples are bagged and numbered either on site at the drill rig or at the on site core handling facility.

Samples sent to laboratories outside Bendigo were sealed in bags in lots of about 10 and sent using commercial freight companies with tracking systems to the relevant laboratories. On arrival at the laboratory, the list of samples sent is matched to the actual samples received and confirmation is sent by either fax or email.

Analytical laboratories have operated in Bendigo during the periods 1992 – 2000 and 2005 to present. During these periods individual samples from the drill rig or core shed have been placed in a container within the mine security gate and collected daily by laboratory staff. Again, on arrival at the laboratory, the list of samples sent is matched to the actual samples received and confirmation is sent by either fax or email.

Work undertaken by employees of the company was limited to core logging and the markup, cutting and bagging of samples. All other sample preparation and analysis was conducted off-site at commercial laboratories.

13.3.2 Data Security

Data security is ensured through the use of an ‘Acquire / SQL Server’ database of all company exploration drilling information. This database includes all assays, geological and geotechnical information. As well as data interrogation, the database allows automated error checking as new data is entered. The database is backed up to tape incrementally daily and fully weekly.

Access to the database is controlled by user login permissions and the Acquire software.

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13.4 Adequacy of Procedures

It is the opinion of the authors that the sample preparation, security and analytical procedures are adequate and have been appropriately applied over the life of the project to ensure that the data is of reasonable quality and representivity.

14. Data Verification

14.1 Database Validation

The exploration drilling carried out by Perseverance at Fosterville has routinely included quality assurance and quality control checks. The nature of these checks has evolved through time and these are described below. In addition, sampling QAQC consultants SMP Consultants reviewed the sampling, analytical and data storage procedures used in the current drilling programme to May 2002 (Crase, 2002). Data systems reviews of the Exploration database were also undertaken by IO Digital Systems in 2004 and 2006 (Kelemen, 2004; McConville, 2006).

The database includes numerous automated data validation methods. The database structure and the use of primary key fields prevents certain types of invalid data (e.g. overlapping sample intervals) from being stored in the database. Also, numerous checks are performed on the data when it is imported (e.g. assay QAQC performance gates, variation in downhole surveys from previous survey).

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Prior to 2000, the geological data was entered directly into the database by hand from the original hardcopy geological log with a manual validation system. Since 2001, all geological data has been downloaded directly from IPAQ hand held logging devices into the database with similar automatic checks as used for the assays. Immediately after the IPAQ is downloaded a hard copy of the geological log is printed to provide an extra back up of the data.

The downhole survey data is the only data hand entered into the database. Allwood (2003) reports a program conducted in 2002 where approximately 10% of the SPD holes were randomly selected for checking the database against the original survey shots. This check found several errors so it was decided to check the entire downhole survey database against the original surveys shots. All errors found were corrected. Underground diamond drill hole (prefix UD) traces are visually checked in MineSight software against the design trace, as soon as the downhole surveys are entered into the database.

14.2 QAQC

14.2.1 Assays

The assay QAQC programme comprises four main strands: the insertion of standards, laboratory duplicates from pulp, laboratory field duplicates and umpire laboratory duplicates. Blank standards are not used because there is a sharp visual grade contrast between mineralisation and waste which provides a natural blank.

All current drilling programmes include the use of four gold mineralised standards provided by Gannet Holdings P/L (ST148, ST109/0285, ST73/7192 and ST43/7194) and one standard prepared from approximately 500kg of Fosterville Sulphide mineralisation from previous RC drilling (AA). One standard is submitted with each drill hole which equates to an effective submission rate of one standard per 30 samples.

The average results for both AMDEL and OSL standard data fall very close to the expected values. All the values falling outside the expected ranges have been investigated and where appropriate, batches are either re-assayed from stored pulps or re-sampled from remaining drillcore.

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Laboratory duplicates of gold are highly repeatable with R² correlation coefficients of 0.99 for both AMDEL and OSL data. There was also a program of inter-laboratory check assays undertaken in 2002 comparing the AMDEL results to two other commercial laboratories – Aminya and Genalysis. The two batches (147 samples) sent to Aminya returned an average of 9% higher with an R² correlation coefficient of 0.993. The Genalysis results were 2% lower with an R² correlation coefficient of 0.996. The inter-laboratory check samples range in grade from below detection (<0.01 g/t Au) to 45 g/t Au. All the inter-laboratory check data is presented in Allwood (2003).

The laboratories provide assay data in digital form as well as hard copy. The digital data is imported directly into the database with a variety of automatic quality control checks preventing sample number mismatches, sample interval overlap, etc.

14.2.2 Surveys

Allwood (2003) details the results of downhole surveys repeated using both an Eastman camera and an EMS tool. The EMS downhole surveys agreed with the single shot surveys to within 0.1° in dip and 2° in azimuth resulting in a total average variation of 0.4 m per 100 m downhole. The repeated Eastman surveys have an average variation of 0.6° in azimuth and 1.6° in dip, reflecting the precision of the Eastman camera survey tool. Comparing the drillhole traces plotted using the Eastman data with the EMS data shows that the variation in hole location due to survey method is considerably less than the variation in hole trace caused by the use of different drillhole de-surveying algorithms.

14.3 Data Verification

In addition to the quality control and data verification procedures discussed in detail above, the Qualified Persons preparing the Mineral Resource estimates have further validated the data upon extraction from the database prior to resource interpolation. This verification used MineSight drill views as the primary tool to identify data problems. This allowed the omission of holes if they were of questionable quality, for example due to low quality sample techniques or incomplete assaying. When coupled with the more mechanical check processes ensuring high quality is entering the database in the first place, these checks were effective in allowing the Qualified Persons to be confident that the data was geologically coherent and of appropriate quality.

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15. Adjacent Properties

As shown in Figure 2, the Fosterville Mine Lease (MIN5404) is completely enveloped by exploration leases held by Northgate (through Perseveration Exploration Pty Ltd). There is Fosterville Fault-related mineralisation identified in the Goornong area (5 kilometres to the north of MIN5404) and the Mills-Hallanan’s area (2 kilometres to the south), however the exploration of these prospects is only at an early stage and not relevant to discuss further in relation to this Technical Report.

16. Mineral Processing and Metallurgical Testing

The following section details the metallurgical testwork conducted on a range of Fosterville ores from 1995 to 2003, with particular focus on the testwork that contributed to the Fosterville Bankable Feasibility Study in 2003 (Persverance, 2003). Multiple batch flotation testwork campaigns and two pilot flotation testwork campaigns were completed for Perseverance by Metallurgy International (MI), Amdel Limited (Amdel), and Ammtec. The biological oxidation technologies offered by both Bactech (Australia) Ltd (Bactech) and Goldfields limited (GFL) were extensively tested and the latter technology was selected.

Samples used for testwork were selected by Fosterville personnel in conjunction with Mr David Foster of Metallurgy International. Effort was made to ensure, where possible, the samples selected for testwork were representative of the respective ore bodies that make up the resource. The major ore bodies Phoenix, Falcon, Ellesmere and Harrier were all subjected to testwork as either major composites or variability samples. The location of the metallurgical testwork samples within the various orebody structures is provided on the plan and sections of the ore bodies and presented in Figure 8.

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Figure 8: Longitudinal projection showing location of the metallurgical testwork conducted from 1996-2003.

Details of the samples collected and the testwork undertaken on each sample is contained in Table 8.

Sample No	Lode	Hole No	Metres			Drill Grade gAu/t	Comments	Sulphide (Y/N)	Weathering (Fresh /Clay)	Purpose of Test
			From	To	=RL
FSVMETS3	Phoenix	SPD106	114.0	139.4	35	4.98	Mineralised shale topped sands	Y	F	Concentrate for oxidation testwork
FSVMETS5	Falcon	CEND113			50 -	9.87	From 1996 JK	Y	C?	Undiluted
		CEND111			100		drop weight			Falcon for BIOX
		CEND089					tests,			to compare
							20 kg			previous BIOX
							CEND113, 14
							kg CEND111,
							22 kg CEND089
FSVMETS6	Falcon	CEND113			50 -	9.07	From 1996 JK	Y	C?	Diluted Falcon
		CEND111			100		drop weight			for BIOX sample
		CEND089					tests,
							20 kg
							CEND113, 14
							kg CEND111,
							22 kg CEND089
FSVMETS8	Ellesmere	SPD077	191.8	200.1	10	5.60	Mineralisation	Y	F	Diluted (~10%)
		SPD027	239.3	242.7	-60		with 0.7 m HW	Y	F	Ellesmere for
		SPD028	266.8	274.2	-100		dilution for	Y	F	BIOX sample
		SPD063	149.3	164.0	20		each hole	Y	F
FSVMETS9	Phoenix -	SPD061	431.6	441.3	-240	6.81	Mineralisation	Y	F	Diluted (~10%)
	Deep	SPD072	476.3	482.0	-300		with 0.7 m HW	Y	F	deep Phoenix
		SPD026A	442 3	452.7	-280		dilution for	Y	F	for BIOX sample
		SPD053	398.8	405.5	-230		each hole	Y	F
		SPD058	417.0	420.7	-230			Y	F
FSVMETS10	Ellesmere	CELD061			50 -	2.82	From 1996 JK	Y	C	Undiluted
		CELD062			100		drop weight			Ellesmere BIOX
		CELD063					tests,			samples
							20 kg from			compare
							each hole			previous BIOX
										work

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Sample No	Lode	Hole No		Metres		Drill Grade gAu/t	Comments	Sulphide (Y/N)	Weathering (Fresh /Clay)	Purpose of Test
			From	To	=RL
FSVMETS11	Falcon	CEND113			50 -	8.71	From 1996 JK	Y	C	Test recovery to
		CEND111			100		drop weight			gravity
		CEND089					tests, 20 kg			concentrate on
							from each hole			graphitic
										mineralisation
FSVMETS14	Phoenix	SPD147	252.3	281.0	-95	2.53	Phoenix	Y	F	Phoenix for
							"ROM" ore			BIOX / POX
							some internal			testwork
							dilution, core
							has ~10% Ag
							removed for
							assay
FSVMETS20	Phoenix	SPD147	305.0	310.3	-130	4.25	High Sb U/G	Y	F	Phoenix for
	BIOX/POX	SPD102	361.7	382.1	-205	22.50	mineralisation			BIOX / POX
		SPD109	200.3	201.3	-25	5.79	to make total			testwork
							1.28 kg for
							doping
							1.94 kg for
							doping
FSVMETS21	Phoenix	SPD148			-65	5.83		Y	F
FSVMETS22	Falcon	CEND113			50 -	9.15	From 1996 JK	Y	C	Shallow Falcon
		CEND111			100		drop weight			for POX
		CEND089					tests,			testwork
							23.4 kg
							CEND111, 20.5
							kg CEND113,
							46.3 kg
							CEND089
FSVMETS23	Ellesmere	CELD061			50 -	2.82	From 1996 JK	Y	C	Shallow
		CELD062			100		drop weight			Ellesmere for
		CELD063					tests,			POX testwork
							13.2 kg
							CELD061, 10.1
							kg CELD062,
							21.1 kg
							CELD063

Table 8: Details of the metallurgical samples collected during 1996-2003 and the testwork undertaken.

16.1 Results

A program of batch flotation testwork (December 2002 to February 2003) was conducted at Ammtec on a range of composites from each of the Phoenix, Falcon and Ellesmere orebodies and on fresh and weathered (clay) ore types. Sample details can be seen in Table 8 above and details of the results are contained in Table 9 below.

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Ammtec Test Number	Fosterville Sample ID	Calculated Head Grade					Cumulative Laboratory Float Time min	Concentrate Mass Pull %	Cumulative Concentrate Assays					Cumulative Recovery to Concentrate
		Au g/t	S %	As g/t	CTOT %	CORG %			Au g/t	S %	As %	CTOT %	CORG %	Au %	S %	As %	C org %
Phoenix Ore
							15	6.73	80.30	29.0	7.0	0.56	0.39	97.80	97.70	94.90	34.22
RG4641A	FSV3 1D	5.53	2.00	5000	0.82	0.08	25	7.84	69.73	25.0	6.1	0.57	0.36	98.77	98.43	96.39	37.14
	Fresh						35	8.59	63.85	22.9	5.6	0.57	0.34	99.08	99.08	96.92	38.83
							45	9.37	58.62	21.0	5.2	0.58	0.32	99.28	99.28	97.27	40.17
							15	8.28	74.00	23.9	8.2	0.88	0.57	90.13	96.04	98.09	44.06
RG4674	FSVMETS9	6.8	2.06	6900	1.04	0.11	25	9.44	70.95	21.1	7.2	0.88	0.54	98.49	96.85	98.09	47.20
	Fresh						35	10.50	63.80	19.0	6.5	0.86	0.50	98.49	97.21	98.44	48.97
							45	11.59	57.80	17.3	5.9	0.84	0.47	98.49	97.42	98.66	50.50
							15	5.39	51.80	16.9	8.7	1.11	0.60	94.23	95.65	95.72	38.26
RG4790	FSVMETS14	2.96	0.95	4900	1.05	0.08	25	7.20	39.39	12.8	6.6	1.09	0.50	95.75	96.87	97.12	42.76
	Fresh						35	8.42	33.85	11.0	5.7	1.07	0.45	96.19	97.32	97.52	44.64
							45	9.67	29.60	9.6	5.0	1.05	0.41	96.49	97.62	97.79	46.54
							15	8.31	58.90	29.6	6.0	0.67	0.54	93.06	95.49	91.08	33.23
RG4681	FSVMETS20	5.26	2.57	5500	1.07	0.13	25	10.01	51.30	25.1	5.4	0.70	0.50	97.70	97.70	97.30	37.30
	Fresh						35	11.71	44.20	21.6	4.6	0.70	0.43	98.30	98.10	98.40	39.70
							45	13.06	39.71	19.4	4.2	0.72	0.43	98.60	98.30	98.40	41.50
							15	7.06	71.50	20.9	10.2	0.89	0.51	96.05	94.22	96.88	37.19
RG4791	FSVMETS21	5.25	1.57	7400	1.2	0.1	25	8.47	60.34	17.6	8.6	0.89	0.47	97.32	95.30	98.03	40.85
	Fresh						35	9.63	53.36	15.6	7.6	0.89	0.43	97.81	95.71	98.42	43.12
							45	10.81	47.70	13.9	6.8	0.89	0.40	98.13	96.01	98.68	44.70
Falcon Ores
							15	9.37	130	28.3	10.6	1.26	1.10	95.06	96.01	95.40	54.90
RG4672	FSVMETS5	12.81	2.76	10400	0.95	0.19	25	11.11	111.54	24.3	9.1	1.19	1.00	96.74	97.69	97.05	59.45
	Clayey						35	12.18	102.17	22.2	8.3	1.15	0.94	97.08	98.03	97.45	61.20
							45	13.13	95.16	20.7	7.8	1.12	0.90	97.52	98.43	97.83	62.98
							15	8.80	124.00	26.7	10.9	1.50	1.34	97.84	96.90	97.45	85.38
RG4686							25
*	FSVMETS6	11.3	2.42	9800	0.17	0.14		10.55	103.41	22.7	9.3	1.43	1.22	97.84	98.81	99.34	93.25
	Clayey						35	11.83	92.21	20.4	8.3	1.37	1.13	97.84	99.53	100.00	97.15
							45	12.90	84.59	18.8	7.6	1.33	1.07	97.84	100.00	100.00	100.00
							15	8.37	119.00	26	11.4	1.21	0.93	96.45	95.84	97.93	81.89
RG4685	FSVMETS11	10.33	2.27	9700	0.14	0.1	25	10.10	100.40	22	9.6	1.19	0.86	98.21	97.89	99.68	91.54
	Clayey						35	11.35	89.35	19.7	8.6	1.16	0.81	98.21	100.00	99.68	96.54
							45	12.42	81.70	18.1	7.9	1.14	0.77	98.21	99.04	100.00	100.00
							15	7.59	128.00	28.7	5.9	1.16	0.93	89.88	94.15	78.59	50.24
RG4782	FSVMETS22	10.81	2.31	5700	0.92	0.14	25	9.17	111.98	24.6	5.6	1.13	0.82	94.93	97.60	89.00	56.62
	Shallow						35
	Sulphide							10.24	101.62	22.2	5.1	1.10	0.87	96.22	98.52	91.45	59.98
							45	11.13	94.17	20.6	4.8	1.07	0.78	96.92	99.04	92.78	62.07
Ellesmere Ores
							15	4.54	54.80	25.6	8.3	1.46	1.21	84.58	95.77	92.25	44.60
RG4673	FSVMETS10	2.88	1.21	4100	0.61	0.12	25	5.58	45.57	21.1	6.9	1.38	1.09	86.36	97.09	93.83	49.56
	Clayey						35	6.67	38.49	17.8	5.8	1.27	0.97	87.19	97.67	94.53	52.65
							45	7.56	34.36	15.7	5.2	1.20	0.90	88.31	98.09	95.05	54.99
							15	5.11	69.50	25.1	5.9	0.89	0.55	94.90	95.74	89.74	33.82
RG4783	FSVMETS23	3.75	1.34	3400	0.65	0.08	25	6.56	55.47	19.9	4.8	0.89	0.51	97.20	97.49	93.62	40.26
	Shallow						35
	Sulphide							7.56	48.43	17.4	4.2	0.88	0.47	97.82	97.97	94.57	42.91
							45	8.36	44.00	15.8	3.8	0.86	0.45	98.16	98.29	95.09	44.91
							15	6.64	70.20	29.1	7.9	1.23	0.93	94.71	95.86	95.57	51.75
RG4675	FSVMETS8	4.92	2.02	5500	1.09	0.12	25	7.78	61.20	25.1	6.9	1.21	0.87	96.75	97.00	97.06	56.44
	Fresh						35	8.93	53.70	22	6.0	1.18	0.80	97.44	97.51	97.68	59.91
							45	10.21	47.42	19.3	5.3	1.13	0.73	98.36	97.77	98.05	62.37

Table 9: Flotation test results

In addition, a number of concentrate samples were generated over the same period for BIOX® and cyanide leach testing. Details of the samples are contained in Table 10 with the BIOX® results reported in Table 11 and the cyanide leach results reported in Table 12.

Composite Identification	Orebody	Ore Type	Drill Hole	Core Type	Intercept	Approx. RL	Comment
FSVMETS 5	Falcon	Clay S	CEND 089 CEND111 CEND 113	PQ PQ PQ	31.0 - 109.5 106.4 - 128.8 91.2 - 109.5	50 - 100 50 - 100 50 - 100	Variability Undiluted
FSVMETS 6	Falcon	Clay S	CEND 089 CEND111 CEND 113	PQ PQ PQ	31.0 - 109.5 106.4 - 128.8 91.2 - 109.5	50 - 100 50 - 100 50 - 100	Variability Diluted
FSVMETS 7	Phoenix	Fresh S	SPD 106	PQ	114.0 - 139.4	35	Rem. METS 3
FSVMETS 8	Ellesmere	Fresh S Fresh S Fresh S Fresh S	SPD 027 SPD 028 SPD 063 SPD 077	¼ NQ2 ¼ NQ2 ¼ NQ2 ¼ NQ2	239.3 - 242.7 266.8 - 274.2 149.3 - 164.0 191.8 - 200.1	-60 -100 20 10	Variability 0.7m Hanging Wall Dilution

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FSVMETS 9	Phoenix	Fresh S	SPD026A SDP053 SPD058 SDPD061 SPD072	¼ NQ2 ¼ NQ2 ¼ NQ2 ¼ NQ2 ¼ NQ2	442.3 - 452.7 398.8 - 405.5 417.0 - 420.7 431.6 - 441.3 476.3 - 482.0	-280 -230 -230 -240 -300	Variability 0.7m Hanging Wall Dilution
FSVMETS 10	Ellesmere	Clay S	CELD 061 CELD 062 CELD 063	PQ PQ PQ	57.0 - 86.0 50.0 - 67.3 24.0 - 62.0	50 - 100 50 - 100 50 - 100	Variability Undiluted Ellesmere
FSVMETS 20	Phoenix	Fresh S	SPD 102 SPD 109 SPD 147	½ NQ2 ½ NQ2 PQ	381.7 - 382.1 200.3 - 201.3 305.0 - 310.3	-205 -25 -130	Variability High Sb BIOX®
FSVMETS 21	Phoenix	Fresh S	SPD 148	PQ	207.0 - 240.0	-65	Variability
FSVMETS 24	Phoenix	Fresh S	SPD 118 SPD 118 SPD 116 SPD 117 SPD 107 SPD 015	¼ NQ2 ¼ NQ2 ¼ NQ2 ¼ NQ2 ¼ NQ2 ¼ NQ2	314.0 - 327.0 236 - 241 176.9-186.5 286.0-307.9 327.0-336.4 243.8-263.0		Variability

Table 10: Summary of samples used for BIOX® and cyanide leach testwork.

			Mass	Reagent Consumption		Sulphide	Arsenic
Sample ID	Orebody / Ore Type	Sample	Gain			Oxidation	Dissolution
			%	Lime	Net Acid	%	%
				kg/t BIOX® Feed	kg/t BIOX® Feed
FSVMETS 5	Clayey Falcon ore	Sample A	34.6	181	-238	95.0	93.1
		Sample B	19.4	201	-302	91.5	94.3
FSVMETS 6	Clayey Falcon ore	Sample A	2.8	137	-196	98.5	95.9
		Sample B	1.1	166	-253	96.9	95.0
FSVMETS 7	Fresh Phoenix ore	Sample A	5.9	137	-200	96.8	93.6
		Sample B	2.8	166	-246	97.6	95.1
FSVMETS 8	Fresh Ellesmere ore	Sample A	18.5	163	-244	99.2	80.5
		Sample B	8.9	129	-179	99.9	89.3
FSVMETS 9	Fresh Phoenix ore	Sample A	23.4	163	-241	98.7	40.5
		Sample B	12.3	150	-211	99.3	46.8
FSVMETS 10	Clayey Ellesmere ore	Sample A	2.6	110	-157	99.9	95.1
		Sample B	2.9	118	-171	99.1	93.6
FSVMETS 20	Fresh Phoenix ore	Sample A	2.1	104	-131	99.2	44.1
		Sample B	32.9	144	-214	98.9	29.1
FSVMETS 21	Fresh Phoenix ore	Sample A	-2.4	85	-112	71.1	45.8
		Sample B	6.6	141	-197	99.9	82.3
FSVMETS 24	Fresh Phoenix ore	Sample A	2.6	118	-161	96.8	53.4
		Sample B	-1.1	111	-152	98.9	59.8
Average			9.8	140	-200	96.6	73.7

Table 11: Summary of BIOX® variability testwork.

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Sample ID	Orebody / Ore Type	Testwork Method	Reagent Consumption		Residual Cyanide CNFREE ppm
			Lime kg/t BIOX® Feed	Net Acid kg/t BIOX® Feed		Gold Extraction %
FSVMETS 5	Clayey Falcon ore	As Received	0.1	11.3	4,119	3.0
		BIOX® Residue A	2.3	12.5	2,393	92.9
		BIOX® Residue B	1.8	12.2	2,930	93.5
FSVMETS 6	Clayey Falcon ore	As Received	0.4	13.0	3,684	8.3
		BIOX® Residue A	2.3	13.8	2,582	94.2
		BIOX® Residue B	1.3	12.4	2,437	92.5
FSVMETS 7	Fresh Phoenix ore	As Received	0.2	9.9	4,467	0.7
		BIOX® Residue A	2.2	14.8	2,379	94.1
		BIOX® Residue B	1.3	12.9	2,698	90.3
FSVMETS 8	Fresh Ellesmere ore	As Received	0.0	27.3	696	4.6
		BIOX® Residue A	5.2	17.4	1,769	91.7
		BIOX® Residue B	4.2	17.5	2,292	93.1
FSVMETS 9	Fresh Phoenix ore	As Received	0.0	28.5	348	2.1
		BIOX® Residue A	23.3	19.0	1,566	95.8
		BIOX® Residue B	9.9	23.3	406	85.2
	Repeat	BIOX® Residue B	12.6	17.8	3,553	94.7
FSVMETS 10	Clayey Ellesmere ore	As Received	0.1	12.7	3,742	37.7
		BIOX® Residue A	2.6	16.0	2,350	95.0
		BIOX® Residue B	1.5	13.4	2,495	93.2
FSVMETS 20	Fresh Phoenix ore	As Received	0	27.8	493	4.6
		BIOX® Residue A	27.2	25.5	1,102	88.5
		BIOX® Residue B	53.2	28.0	1,944	91.7
FSVMETS 21	Fresh Phoenix ore	As Received	0	25.9	1,015	1.7
		BIOX® Residue A	6.4	13.8	3,626	95.3
		BIOX® Residue B	6. 9	15.4	3,481	94.0
FSVMETS 24	Fresh Phoenix ore	As Received	0.02	21.5	841	8.6
		BIOX® Residue A	13.9	23.7	1,653	89.3
		BIOX® Residue B	10.9	20.2	2,089	89.7
Average		BIOX® Only	9.9	17.0	2,408	92.7

Table 12: Summary of cyanide leach of BIOX® variability residues.

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17. Mineral Resource and Mineral Reserve Estimates

Table 13 below outlines the Mineral Resources for the Fosterville Gold Mine as at December 31, 2007.

Mineral Resources - Fosterville
Classification		Measured			Indicated			Inferred
		Tonnes (kt)	Grade (g/t Au)	Insitu Gold (kOz)	Tonnes (kt)	Grade (g/t Au)	Insitu Gold (kOz)	Tonnes (kt)	Grade (g/t Au)	Insitu Gold (kOz)
Fosterville Fault Zone Sulphide Resources
Central Area	Upper	2,326	2.45	183	404	1.87	24	146	2.23	11
	Lower	703	4.71	107	5,766	4.74	879	6,108	3.58	703
Southern	Upper	152	2.04	10	507	2.32	38	1,166	1.93	72
Area	Lower							5,648	3.34	607
Northern	Upper	333	1.83	20	240	1.23	10	25	0.99	1
Area	Lower
Robbin's Hill Area Sulphide Resources
	Upper	39	1.19	1	2,303	1.31	97	1,398	1.44	65
Combined
	Lower				296	2.59	25	853	2.53	69
Sulphide Upper		2,849	2.34	214	3,454	1.52	169	2,735	1.69	148
Sulphide Lower		703	4.71	107	6,062	4.64	903	12,609	3.40	1,380
Total Sulphide		3,553	2.81	321	9,516	3.50	1,072	15,344	3.10	1,529
Total Oxide		606	1.17	23	1,857	1.10	66	598	1.36	26
Total Oxide & Sulphide		4,159	2.57	343	11,372	3.11	1,138	15,942	3.03	1,555

Table 13: Mineral Resources for Fosterville Gold Mine as at December 31, 2007.

Notes:

1.

For the Mineral Resources estimate, the Qualified Person for the Central Area is Ian Holland and for all other areas is Simon Hitchman. Their details and qualifications can be seen in Section 23 of this report

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.

2.

The Mineral Resources for the Central Area reported in Table 1 are inclusive of the Mineral Reserves for the same area (reported in Table 2 below).

3.

Cut-off grades applied to Mineral Resources are 0.5 g/t Au and 0.7 g/t Au for oxide and sulphide mineralisation respectively above 5050mRL (approximately 100 metres below surface) which is deemed to be potentially open-pittable. The Mineral Resource cut-off grade applied for material below this point is 2.0 g/t Au.

4.

Mineral Resources have been rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summation may occur due to rounding.

Mineral Reserves - Fosterville
Classification		Proven			Probable		Total
						Contained
	Tonnes	Grade	Contained Gold	Tonnes	Grade	Gold	Tonnes	Grade	Contained Gold
	(kt)	(g/t Au)	(kOz)	(kt)	(g/t Au)	(kOz)	(kt)	(g/t Au)	(kOz)
Central Area
Phoenix	49	3.45	5	4,444	4.48	640	4,493	4.47	646
Falcon	87	4.00	11	95	3.69	11	182	3.84	22
Ellesmere				550	4.61	82	550	4.61	82
Kink				390	3.79	48	390	3.79	48
Total	136	3.80	17	5,480	4.43	781	5,616	4.41	798

Table 14: Mineral Reserves for Fosterville Gold Mine as at December 31, 2007.

Notes:

1. For the Mineral Reserves estimate, the Qualified Person is Brad Evans. His details and qualifications can be seen in Section 23 of this report.

2. The Mineral Reserve estimate used a gold price of AUD$750/ounce. The cut-off grades applied ranged from 1.2 g/t to 3.6 g/t Au for underground sulphide ore depending upon width, mining method and ground conditions.

3. Dilution of 5-30% and mining recovery of 70-95% were applied to the Mineral Reserves dependant upon mining method.

4.Mineral Reserves have been rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summation may occur due to rounding.

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The Mineral Resources and Mineral Reserves reported above are largely the result of work undertaken in June 2007 and reported by Perseverance under Australian reporting requirements in accordance with the JORC Code. Although there have been some minor updates, the majority of the models remain unchanged from this period aside from being depleted for material mined from July 1, 2007 to December 31, 2007. Where there has been new data acquired (drilling or otherwise) in this period, it is the opinion of the authors that this new data has not made a material difference to this estimate.

In all cases, the Qualified Persons have reconciled the estimates to CIM standards as prescribed by NI 43-101.

For discussion on other issues that may impact upon the Mineral Resource and Mineral Reserve estimates such as taxation, marketing, legal and environmental considerations, the reader is referred to Section 24 of this report.

The location and extents of the block models for each of these areas are displayed in Figure 9 below. Current mining activities are confined to the Central Area (Falcon, Ellesmere and Harrier models). This area will be described in detail below and will be followed by a description of the other areas (Northern, Southern and Robbins Hill models).

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Figure 9: Plan showing mining lease and the area covered by each of the block models.

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17.1 Central Area

The Central Area is defined as the zone between 6600mN and 10000mN (Mine Grid). The controlling features in this area are the Fosterville Fault and a number of footwall splay structures which control mineralisation (see Figure 10 for a representative cross section). The Mineral Resource and Mineral Reserve modelling process for this area is described below.

Figure 10: Cross-section though Phoenix and Ellesmere Shoots showing mineralization domains, faults, bedding formlines and the location of hanging wall black shale.

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17.1.1 Geological Models

In order to constrain the resource models, a number of three-dimensional geological models were generated for each ore zone using Minesight software. The models produced were of three types:

 1.

structural wireframe models

2.

mineralisation wireframe models

3.

waste wireframe models

The structural models contained three-dimensional wireframe surfaces of major faults and splay structures as identified by mapping in the open pits and underground, and from diamond drilling. The mineralisation model defined the interpreted gold-bearing mineralised zones and was constrained either by structural, lithological or grade boundaries. The waste model is a 10-15 metre envelope surrounding the mineralisation model (see Figure 11).

Figure 11: Section showing elements used in the construction of mineralisation domain wireframes.

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A string outlining the natural limit to mineralisation (generally 0.3 g/t Au to 1.0 g/t Au) is created for each drill section. This means that the interpretations are completed on 6.25 metre sections in areas of open pit grade control drilling and on 25 metre, 50 metre and 100 metre sections where there is only surface exploration drilling.

The string is ‘snapped’ to an assay boundary so that an assay is either entirely within the mineralisation interpretation or entirely excluded from the mineralisation interpretation. The mineralisation strings are then linked in 3-dimensions to create 3D mineralisation envelopes. Separate mineralisation envelopes were created for geologically distinct zones.

Information from mapping of the sill development in the Phoenix and Falcon ore zones was also used during the modelling process, as were digital photographs of the diamond drill core. This data was used to confirm interpretations of mineralisation boundaries and significant fault zones.

17.1.2 Domaining

Based on observed variations in geology, variography, geochemistry, statistics and spacial location within the Fosterville mine area, mineralisation in the Central Area was divided into four distinct domains. The domains corresponded to:

1.

Falcon Ore Zone

2.

Phoenix Ore Zone

3.

Ellesmere Ore Zone

4.

Harrier and Splays Ore Zone

The domains are differentiated primarily on structural setting and mineralisation plunge. The host geology of the mineralisation is consistent across all domains and is made up of inter-bedded sandstones and shales. Additionally, the mineralisation forms on either the hanging wall or footwall to the main structures, depending on the bedding orientations. Surrounding all the ore domains is a waste domain and this was used to generate the waste gold grades in the immediate vicinity of the mineralisation.

The Falcon and Ellesmere Ore Zones are located on the footwall to the Fosterville Fault and separate domains based on mineralisation plunge either side of the culmination. The Phoenix Ore Zone is located on the hangingwall of the Phoenix Fault and is in a distinctly different structural setting than the Falcon and Ellesmere domains. The Harrier and Splay Ore Zones are located along minor structures between the Phoenix and Fosterville Faults and the domain is made up of numerous localised ore pods. Each of these domains clearly distinctly different gold grade distributions.

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17.1.2.1 Falcon Domain

The Falcon domain encompasses all the mineralisation north of 8900mN in the Falcon FF, Vanessa Splay 1, Vanessa Splay 2 and the Vanessa Splay 3 mineralisation envelopes. A plunge reversal occurs between 8800mN and 8900mN and all of the mineralisation between 8900mN and 11000mN plunges gently to the north. The vast majority of the mineralisation in the Falcon domain occurs on the Fosterville Fault and dips about 70° to the west. Most of this domain is relatively shallow (less than 150 metres below surface) and has been drilled by either RC drilling grade control drilling on 6.25 metre sections or by RC and diamond exploration drilling on 20 metre sections. Assays, composites and blocks occurring within the Falcon domain are coded so that DOMA=1.

17.1.2.2 Phoenix Domain

The second domain is the Phoenix domain which includes all the Phoenix main and Phoenix mid-splay mineralisation envelopes. The mineralisation in the Phoenix domain plunges 15° to 20° to the south. The mineralisation on the Phoenix Fault is very consistent in width and geometry. The Phoenix mineralisation dips 45° to 65° to the west. The Phoenix includes a 20 metres wide, 50 to 100 metre high zone extending from 8000mN to 8400mN. The Phoenix domain is open down plunge to the south, but is limited to the north at 8900mN where mineralisation steps from the Phoenix fault on to the Fosterville Fault. Assays, composites and blocks occurring within the Phoenix domain are coded so that DOMA=2.

17.1.2.3 Ellesmere Domain

The Ellesmere domain is the same as the Ellesmere Fosterville Fault mineralisation envelope. The plunge of the mineralisation in the Ellesmere FF domain appears to be 20° to 40° to the south with smaller scale sub-shoots plunging 70° to the west. The dip is about 70° to the west. Assays, composites and blocks occurring within the Ellesmere FF domain are coded so that DOMA=3.

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17.1.2.4 Harrier & Splays Domain

The Splays domain includes the Ellesmere Splay, Kite Splay, Hawk Splay, Harrier Splay and John’s Splay mineralisation envelopes. Mineralisation in the Splays domain dips 55° to 70° to the west and plunges 25° to 35° to the south. Assays, composites and blocks occurring within the Splays domain are coded so that DOMA=4.

Figure 12: Long projection of mineralisation domains in the Central Area.

17.1.3 Drilling Data

After the geological models and subsequent mineralisation domains have been defined, the drill hole assay data used to produce the model was subjected to a number of data preparation processes:

1.

Files containing all drill hole logging and assay data were imported from the Acquire production and exploration databases into MineSight using an automated script.

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2.

The script also codes the drillholes with the appropriate properties from the geological models and constructs a drillhole assay composite file for values inside the mineralisation wireframes.

3.

The files were loaded into MineSight and viewed in order to identify holes that contained obvious erroneous data missed during the validation process. Data that was considered erroneous was either corrected or deleted from the data set. Note: step 1 and 2 were also done prior to the geological models being finalised to ensure the interpretations were finalised on a validated drill hole file.

In combination, the drill hole files used for the Central Area models contained a total of 2,723 drillholes between them to estimate sulphide mineralisation, of which 2,128 (78%) are RC, 354 (13%) diamond core, and 241 (9%) underground face channels.

17.1.3.1 Compositing

All assays falling within the mineralisation envelopes were composited to 2 metres starting from the point at which the drillhole enters the mineralisation envelope. If the last composite is less than or equal to 2 metres in length it is left as is. If a drillhole intersects more than one mineralisation envelope then the compositing starts again as each envelope is entered (see Figure 13). Table 15 below shows the composite statistics.

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Figure 13: Downhole compositing where the width of the domain object is not a multiple of two.

Composite length	N	% of comps	mean length (m)	mean grade (g/t Au)
<1 m	414	2%	0.52m	3.42g/t
1m= and <2m	826	4%	1.31m	3.25g/t
2m	17,856	94%	2m	3.99g/t
Total	19,096	100.0	1.94m	3.95g/t

Table 15: Composite statistics by composite length.

Two metre composites were chosen to reflect the anticipated minimum mining width, to allow across strike variability to be maintained within the data and because the vast majority of RC drilling samples are two metres long. Any composites less than 1m are not used in estimation.

In the compositing process, core loss is treated as unsampled core. This means that a two metre downhole distance becomes a 1.8 metre long composite if there is 0.2 metres of core loss. The average core recovery for all intervals within the domains was 98.3%. For both waste and mineralisation combined, the core recovery was 98.5%.

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17.1.3.2 Variography

In general, the structure of the variograms increases from antimony (pure nugget) through fresh gold, oxide gold, arsenic and sulphur to iron.

In all domains and for all elements except antimony there is a low relative nugget effect (15% to 50%), a short range structure (2 to 10 metres) comprising most of the variance with the remaining variance controlled by a long range structure (20 to 100 metres). The low nugget effect reflects the fine grained, disseminated nature of the sulphide minerals. The short range structures are most likely caused by variation in host lithology due to stratigraphic variation.

Two metres is a typical bedding width and also the composite length, thus it is not possible to resolve structures at a smaller scale. Previous studies show that Au, S, Fe and As values are strongly controlled by lithology being higher in shales and lower in sandstones. The long range structures approximate the spacing of splay faults and so probably reflect high grade zones occurring where splay faults intersect the Fosterville and Phoenix Faults.

Sulphur occurs as disseminated pyrite and arsenopyrite and as vein hosted stibnite. The depositional mechanisms for sulphur are both diagenetic (pyrite) and hydrothermal (pyrite, arsenopyrite and stibnite). The varying occurrence modes and depositional mechanisms of sulphur has resulted in sulphur occurring at percent levels in all rock types in the domains resulting in well structured variograms.

Arsenic occurs as arsenopyrite, precipitated from hydrothermal solutions. The precipitation of hydrothermal arsenopyrite has been controlled at metre to 100 metre scales by proximity to major fluid conduits (faults) and at 10 cm scale by lithology. The variability in arsenopyrite deposition has resulted in less structured variograms compared to the sulphur variograms.

Oxide gold occurs as very finely disseminated free gold grains associated with iron oxides formed as a result of weathering of auriferous sulphide minerals (see fresh gold below). Thus, the large scale controls on oxide gold mineralisation are the same as those controlling fresh gold mineralisation except that there has been some smoothing and smearing of grade variability as a result of gold remobilisation during weathering. Consequently, the oxide gold variograms are somewhat more structured than the fresh gold variograms and have longer ranges, albeit with slightly higher nugget effects.

66

Fresh gold predominantly occurs in arsenopyrite and so is subject to the same factors that control the variability of arsenic in space. In addition, the gold content of the arsenopyrite is highly variable and so the fresh gold variograms are the least structured. The relative nugget effect is 16% in the Falcon domain, 24% in the Phoenix domain, 39% in the Ellesmere FF domain and 47% in the Splays domain. The higher nugget effect in the Splays domain is probably because the Splays domain is a composite of mineralisation hosted on several poorly drilled splay faults. As the drilling density increases on these splay faults, further subdivision of the splays domain will be possible and this should result in better structured variograms with lower nugget effects.

17.1.4 Resource Modelling

17.1.4.1 Block Models

For reasons of data handling, the Central Area was divided into 3 separate block models with the following extents and block dimensions:

Table 16: Central Area block model dimensions.

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All models use Ordinary Kriging to interpolate grades and there are no top cuts applied.

17.1.4.2 Search Criteria

Gold, arsenic, sulphur, iron and antimony grades are only interpolated into blocks meeting the following criteria:

1.

Greater than 1% of the block volume is inside one of the domain envelopes

2.

Blocks within one of the resource category solids

3.

 Blocks whose search ellipse includes at least 5 composites

The search ellipse geometries were chosen to reflect the geology, variogram model and drill spacing of the relevant zone so that a block could ‘see’ at least the nearest sections along strike and holes up or down dip.

Similarly, only composites meeting the following criteria are used to interpolate any one block:

1.

Composites (to a maximum of 15 composites) must be within within the search ellipse dimensions and search area limits.

2.

Where more than 15 composites lie within the search ellipse, the 15 closest composites in anisotropic ellipsoid space are used.

3.

Maximum of 6 composites are used from any split-quadrant of the search ellipse (a split-quadrant is 1/8th of the search ellipse divided in the planes of the major, intermediate and minor ellipse axes).

4.

The domain codes of both the composite and the block must match (ie only composites from within the same mineralisation envelope and the same oxidation material are used to interpolate a block)

5.

Greater than or equal to one metre in length.

To ensure the suitability of the search ellipses used, a search ellipse was created in MineSight for each run file allowing visual inspection of the composites used and kriging weights calculated for the block at the centre of the ellipse (see Figure 14 for examples).

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Figure 14: Search ellipses for Domains 1 (Falcon) and 2 (Phoenix).

The resultant block models are therefore tightly constrained by wireframe models derived from detailed geological interpretation and modelling of the ore zones. This provided the vital basic geological control over the computer-generated grade estimations. A section through the block model is included in Figure 15 below.

Figure 15: Section through the Falcon Block Model (8600mN), showing the Phoenix and Ellesmere mineralisation envelopes, and ore and waste grades.

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17.1.4.3 Bulk Density

The bulk density profile established for the Central Area was the result of considerable testwork undertaken from drill core, primarily using the core volume method. Alwood (2003) outlines a program of 361 bulk density and 580 specific gravity determinations. The specific gravity measurements are from water immersion wax coated core (12% of data) and water immersion of uncoated core. There is no discernible difference in the specific gravity values of the wax coated core and uncoated core.

Data was compared by ore / waste domain and RL. The only significant variation in the bulk density data is an increase with depth. The increase in bulk density with depth is thought to be a result of subtle weathering effects below the base of oxidation causing carbonate dissolution and clay alteration of feldspars. The bulk density shows no variation between ore and waste. This is probably because although the ore has a higher bulk density than the waste, it also has a higher fracture porosity, the two effects cancelling each other out.

The bulk densities used in the resource model (see Table 17 below) reflect the different material types and the increase in bulk density with depth.

RL Definition (m)	Bulk Density (kg/m3)	Description
Surface to base of alluvium	1600	Alluvium
Surface or base of alluvium to base of oxidation	2400	Oxide rock
Base of oxidation – 5050 RL	2560	Clay alteration, carbonate depletion
5050 RL – 5000 RL	2640	Weak clay alteration, carbonate depletion
Below 5000 RL	2720	Fresh

Table 17: Bulk density values used in the resource model.

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17.1.5 Mineral Resource Classification

The mineral resource estimates were generally classified according to the following parameters:

1.

Areas that have underground development directly above and below were classified as Measured Mineral Resources.

2.

Areas drilled up to a spacing of 50 metres x 50 metres were classified as Indicated Mineral Resources.

3.

Areas drilled to a spacing wider than 50 metres x 50 metres were classified as Inferred Mineral Resources.

These parameters may vary subject to the level of geological confidence in specific areas.

Figure 16: Longitundinal section showing resource classification for Falcon and Phoenix ore zones
(ore drives and stope shapes only) - Measured (blue), Indicated (green) and Inferred (red).

17.1.6 Mineral Reserve Estimate

The following sections outline the process undertaken to produce Mineral Reserve estimates from the available Mineral Resource. The section contains descriptions of reserve block design parameters, recovery and unplanned dilution factors, cut-off grades and depletion for mined material. This section of the report has been compiled by Brad Evans who is the Qualified Person for the Mineral Reserve estimate (details in Section 23). Significant input into the process was contributed by Phil Bremner, a consultant working for Mining One.

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17.1.6.1 Mineral Reserve Block Design

The initial stage of the reserve estimation process was the design of the reserve blocks. The mining methods that were considered for the reserve estimation process were sill driving, uphole open stoping, downhole open stoping, modified longitudinal sublevel caving and transverse open stoping. These methods were selected based upon previous experience at the Fosterville mine or because they were considered suitable for the ore zone geometry and geotechnical conditions present and expected.

17.1.6.2 Open Stope Design and Reserve Parameters

Open stope reserve blocks were created to cover the majority of the resource where this was considered to be the most appropriate mining method (see Figure 18 and Table 18 below for coverage of open stoping area). These stope blocks did not necessarily reflect the final stope strike and crown pillar dimensions. Stoping widths vary from 2.5-3 metres out to 20 metres with transverse access.

The open stope reserve wireframe design parameters applied were:

1.

Strike length of 6.25 metres, actual length dictated by grade distribution in block model.

2.

Minimum horizontal bench width of 2.5 metres.

3.

Maximum benching height of 20 metres vertical from backs to floor.

4.

Internal waste incorporated within the stope block design.

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Figure 17: An example of an open stope reserve wireframe design

Mining recovery from open stopes at Fosterville is principally influenced by the following factors:

1.

Accuracy of the geological interpretation.

2.

Accuracy of the production hole drilling.

3.

Stope dimensions.

4.

Sill drive dimensions and position relative to bench stope.

5.

Presence or absence of adjacent filled voids and pillars.

6.

Geotechnical integrity of stope and sill drive walls.

The above factors manifest themselves as ore loss in the following ways:

1.

The need for planned pillars due to accessing of ore blocks (i.e. top down mining sequence).

2.

The need to fill unstable stope voids still containing broken ore stocks.

3.

Frozen rings due to ground movement.

4.

Ore stuck to the Cemented Rock Fill (CRF) of adjacent filled stopes.

5.

Bridged stopes.

6.

Failure of the stope to break back to a main structural plane of weakness.

7.

Unplanned ore pillars left to improve ground support.

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Unplanned dilution in open stopes at Fosterville is a function of the following factors:

1.

Regional geotechnical conditions.

2.

Location of sill drives relative to the open stope.

3.

Width of sill drives relative to the open stope width.

4.

Production drilling accuracy.

5.

 Quantity, quality and type of ground support in sill drive walls.

6.

Speed of ore extraction from active stopes.

7.

Length of time sill drives have been open before stoping commences.

8.

Adjacent filled stopes (either CRF or mullock fill).

In order to correctly apply recovery and dilution factors to all stope blocks in the Mineral Reserve, the mine was subdivided into areas of similar geology, geotechnical conditions and scheduled access. Factors such as orebody dip, rock RQD and development and stope sequence were considered. The current life of mine schedule shows that Phoenix Blocks 2 to 5 will be accessed bottom up as opposed to Block 1 which was locked in to a top down sequence with crown and rib pillars. Table 18 and Figure 18 below shows the recovery and dilution factors that were applied to the reserve blocks:

Code	Description	Recovery Factor -Tonnes	Dilution Factor - Tonnes	Comments
BK1	Phoenix Ore Zone above 4940mRL, Nth of 8250mN	70%	10%	Top down, crown and rib pillars
BK2	Phoenix Ore Zone btw 4745 and 4940mRL, Nth of 7700mN Phoenix Ore Zone above 4940mRL, btw 8100 and 8250mN	90%	10%	Bottom up, primary and secondary stopes
BK3	Phoenix Ore Zone above 4660mRL, btw 7625 and 7875mN	90%	25%	Stope widths between 2.5 and 4m wide
BK4		80%	15%	Stope angles < 50 degrees and narrow
BK5	Phoenix Lower	90%	10%
BK6	Sill Drives	95%	5%
BK7	Ellesmere Block 1	70%	30%	Caving
BK8	Kink	70%	20%	Bottom up, poor ground conditions
BK9	Ellesmere Block 2	90%	10%	Bottom up
BK10	Falcon	70%	30%	Caving

Table 18: Recovery and dilution factors for the reserve blocks displayed in Figure 18 below.

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Figure 18: Longitundinal projection displaying the mining blocks referred to in Table 18 above.

The following formulae were used to apply reserve estimation factors to each reserve block. After application of the reserve estimation factors, each block was assessed as to whether it was to be included in the ore reserve estimate. The final decision on inclusion in the Mineral Reserve inventory was dictated by the spatial location of the block and how the reserve grade compared to the estimated cut-off grade (see Section 17.1.6.4 below).

Open Stope Reserve Tonnes Estimation Formula

Reserve tonnes = (resource tonnes x recovery factor) + (resource tonnes x recovery factor x unplanned dilution factor) tonnes

Open Stope Reserve Ounces Estimation Formula

Reserve ounces = (resource block ounces x recovery factor) +((unplanned dilution tonnes x unplanned dilution grade)/31.1034) oz

Open Stope Reserve Grade Estimation Formula

Reserve grade = (reserve ounces x 31.1034)/ reserve tonnes

17.1.6.3 Modified Longitudinal Sublevel Caving Design and Reserve Parameters

This mining method has been adopted to allow full extraction of the Falcon orebody and the upper section of the Ellesmere orebody with minimal hangingwall exposure. Due to the poor ground conditions associated with the Fosterville Fault and high native carbon content, open stope strike length would have been less than 10 metres and crown pillars would have been required on all levels due to the top down access. This would have lead to up to 40% of the Falcon orebody being left in situ, unrecoverable, in crown and rib pillars including a 15 metre crown pillar beneath the Falcon Open Pit.

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The Modified Longitudinal Sublevel Caving reserve wireframe design parameters used were:

1.

 Strike length of entire Falcon orebody.

2.

Stope width equivalent to mineralisation width.

3.

Maximum stope height of 30m underneath Falcon Open Pit

4.

Internal waste incorporated within the block design.

Mining recovery of the Falcon orebody is dependent on the level of the cave. Overall the average recovery has been estimated at 70%.

An average unplanned dilution factor of 30% was estimated for Falcon due to the blasted ore being bogged from a drawpoint that was ‘choke’ fired against waste from the backfilled open pit. Reserve conversion formulae for Modified Longitudinal Sublevel Caving, and the principles governing whether the resulting block was retained in the reserve total, were the same as for open stope blocks.

17.1.6.4 Cut-off Grades

The table below shows the assumptions used to calculate the average stope and development cut-off grades. Cost assumptions are based on the life of mine plan.

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Gold Price	$/oz	$	750.00
	$/g	$	24.11
Mining cost ($/t)	Falcon	$	37.91
	Ellesmere/Kink	$	47.94
	Phoenix	$	37.91
Non mining cost ($/t)	Processing	$	17.10
	Maintenance	$	5.80
	Sustainability	$	1.45
	Administration	$	3.97
	Total Non-mining cost	$	28.32
Total cost ($/t)	Falcon	$	66.23
	Ellesmere/Kink	$	76.26
	Phoenix	$	66.23
Mill Recovery	Falcon		85%
	Ellesmere/Kink		88%
	Phoenix		88%
Mined Cut-off Grade (g/t)	Falcon		3.23
	Ellesmere/Kink		3.59
	Phoenix		3.12
Development COG (g/t)			1.24
Stoping incremental COG (g/t)			2.00

Table 19: Cut-off grade assumptions used, AUD$ denomination.

For stope production, the cut-off-grade calculation used was:

Cut-off Grade (g/t) = (Total Cost ($/t)) / ( $/gram gold price x Mill Recovery ($/g))

For certain other situations, a lower cut-off grade is applied. For development which is justified for other reasons (ie. access to a higher grade block or infrastructure considerations), the cut-off grade is lowered to reflect that the material only has to cover the non-mining costs to break even. This is only applied if the development material had to be trucked to surface anyway and that it is not displacing higher-grade ore from the mill. Likewise for incremental stoping production where the development has already been mined (ie. for access to a higher-grade block), the cut-off grade is lowered to reflect that the development cost has already been incurred.

17.1.7 Depletion and Results

As mentioned above in section 17, the Mineral Resources and Mineral Reserves reported above are largely the result of work undertaken in June 2007 and reported by Perseverance under Australian reporting requirements in accordance with the JORC Code. Although there have been some minor updates, the majority of the models remain unchanged from this period. The models have been depleted for material mined from July 1, 2007 to December 31, 2007. The process followed involved the generation of surveyed volumes for the mined development and stope areas in MineSight and then physically depleting them from the remaining stope shapes and the resource domains. Allowance is also made in this process for any associated sterilisation that may have occurred.

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Results for the Mineral Resource and Mineral Reserves contained in the Central Area are provided in Table 7 and Table 8 respectively.

17.2 Northern Area

Other areas modelled within the Fosterville Mine Lease include the Northern Area, Southern Area, and Robbins Hill Area. Much of the data collection and systems developed for the Central Area were applied in these areas so for brevity, the ensuing discussions on each area will highlight the geology of each area and process differences from the Central Area modelling.

The northern model is located to the north of the Central model area and is defined as the zone between 10,000mN and 11,500mN (as shown in Figure 9). The controlling structural features from west to east include: the moderately west dipping Hunt’s Fault, several footwall splays and the Fosterville Fault. The geology of the area was assessed by Fosterville staff, later reviewed by Stephen King (2007) and resource modelling undertaken by Kerrin Allwood (2008). Information from the latter two provides the basis for what follows.

17.2.1 Area Geology

The key to mineralization in the Hunt’s Pit area appears to be the relationship between the Hunt’s fault, bedding and a set of splays that strike obliquely to the fault (see Figure 19). Hunt’s fault dips steeply to the west and is bedding parallel on the hanging wall. Bedding in the footwall is oblique to Hunt’s fault and dips to the east where it is intersected by the splay faults. Mineralisation of interest is found in a zone approximately 5-10 metres wide below the Hunt’s fault. It is interpreted that the vertical extent of the mineralisation is controlled by shallower west dipping splays that intersect the Hunt’s fault. The exact position of the intersection of these faults and Hunt’s fault is uncertain. It is known from drill holes that mineralisation exists immediately under Hunt’s Pit, but deeper drilling has failed to intersect mineralisation under the pit. It would appear that mineralisation does not extend more than 20 metres below the pit. There is a lack of diamond holes in this area and the position of major splays is often interpreted upon the data (largely logged quartz content and gold and arsenic assays) gained from RC holes. The diamond holes are the deeper holes, and these intersect the splays near their keel positions and below the mineralisation. It is not clear whether there is one dominant, continuous splay fault striking sub-parallel to the Hunt’s Fault or several that may strike obliquely to the Hunt’s Fault.

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It is also interpreted that there are easterly dipping, bedding parallel capping faults that run between Hunt’s fault and the splay faults. The position of these faults is uncertain as most holes are drilled parallel to these structures. East dipping quartz veins and laminated quartz veins are visible in the northern face of the Fosterville Pit but the continuity of any individual structures along strike is not known.

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Figure 19: Cross section 10,900mN through Hunt’s pit showing the relationship between the Hunt’s Fault, bedding and the set of splays that strike obliquely to the fault.

17.2.2 Geological Models

Constraints to resource models were imposed by three-dimensional geological modelling in MineSight. These included structural wireframes (faults, fold axial planes) and mineralisation wireframes. Wireframes for waste were not created.

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17.2.3 Domaining

Domaining of the Northern Model is simplified as only one gold domain was required. However, owing to geology, weathering and variations in backfill material into the Fosterville oxide pit, the following domains were created and are further discussed:

1.

Gold Domain

2.

Oxidation Domains (Fresh, Transitional, Oxides and Alluvium)

3.

 Back Fill Domain

Only one gold domain was created, although sub-divided by the base of complete oxidation for variography and kriging. This gold domain incorporates mineralisation on the Fosterville Fault, Hunt’s Fault and various splay and linking faults. It was decided to only use one gold domain because the geology of the gold mineralisation on both the Fosterville and Hunt’s Faults is very similar with sub-parallel mineralisation shoots and similar univariate gold statistics. The various splay and linking faults are too small to form individual domains.

For the oxidation domains, four ‘material’ domains were constructed, although only gold mineralisation was interpolated in three of these. The four domains are:

1.

 Fresh (sulphide minerals completely un-oxidised)

2.

Transition (sulphide minerals may be partially oxidised, includes zones of mixed fresh and oxide)

3.

Oxide (sulphide minerals completely oxidised, Fe-carbonates largely oxidised)

4.

Alluvium (near surface transported material, generally barren of gold, largely clay, free digging)

Two surfaces (base of complete oxidation and base of alluvium) were interpreted as strings on each section and then ‘triangulated’ to form surfaces. The base of transition was taken to be 5105mRL because there is insufficient logging of the base of transition to allow a reasonable interpretation of this surface. 5105mRL was chosen as it is 15 metres below the typical base of complete oxidation RL of 5120m and approximates the base of transition where logged. 15 metres is the typical depth of the transition zone in pits that have been mined to date. These surfaces were then intersected to form solid domain wireframes.

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The fresh domain is that volume between the base of transition and the base of the model. The transition domain is that volume between the base of complete oxidation and the base of transition. The oxide domain is that volume between the base of alluvium and the base of oxidation. The alluvium domain is that volume between the pre-mining topography and the base of alluvium.

A back fill domain wireframe was created over the southern part of the Fosterville oxide pit that has been backfilled. The current surface of the top of the back fill has not been surveyed.

17.2.4 Drilling Data

The quality of the drilling is variable in the Northern Area and includes:

1.

5 RAB – Rotary Air Blast,

2.

338 Reverse circulation – Cross over hammer and face sampling hammer variants, and

3.

 38 Diamond core – HQ and NQ2, often with RC pre-collars.

Nine drill holes were omitted owing to uncertain coordinates, dubious down hole surveys, grade or geological mismatch and low quality sample techniques (RAB). Visual checks in MineSight were the primary tool used to identify data grade and geological mismatches.

Subsequent to the drill data review process, assay data were:

1.

 Imported from the acQuire Exploration databases into MineSight using MineSight customizable parameter screens.

2.

 Coded for mineralisation using 3-D gold wireframe solids.

17.2.4.1 Compositing and Coding

Similar to the Central Area, coded northern drill data was composited to 2 metre lengths starting from the point at which the drillhole enters the mineralisation envelope. If the final composite was less than 1.0 metres it was added to the previous composite making a composite with length between 2.0 and 3.0 metres. Final composites between 1.0 and 2.0 metres in length were left as is.

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17.2.4.2 Variography

Experimental normal variograms created for composited and domained gold assay data are generally well structured. The relative nugget effect values are typical for the Fosterville Goldfield (20% to 50%) and, with the exception of the major axis of the fresh model, most of the variance occurs within 10 to 15 metres. The variogram models closely follow the expected geological controls with flat to shallowly north plunging shoots in a steeply west dipping faults.

17.2.5 Resource Modelling

17.2.5.1 Block Models

The Northern Block Model was created to allow modelling of mineralisation north of 10,000mN. XYZ Block dimensions of 4 metres (east) by 10 metres (north) by 5 metres (RL) were used. This block size was chosen after consideration of the maximum drilling density (~25 by 15 metres), variogram model ranges (10 to 20 metres), mineralisation geometry (typical mineralisation width of 4 to 8 metres) and probable mining methods.

17.2.5.2 Search Criteria

Gold and sulphur grades are only interpolated into blocks meeting the following criteria:

1.

Greater than 1% of the block volume is inside one of the domain envelopes.

2.

Blocks whose search ellipse includes at least 4 composites

3.

Blocks whose MATL code is set to Oxide (3), Transitional (2) or Fresh (1).

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Similarly, only composites meeting the following criteria are used to interpolate any one block:

1.

all composites (to a maximum of 30 composites) within the search ellipse dimensions and search area limits outlined in the table below

2.

where more than 30 composites lie within the search ellipse, the 30 closest composites in ellipsoid space are used

3.

 maximum of 5 composites are used from any split-quadrant of the search ellipse (a split-quadrant is 1/8^th of the search ellipse divided in the major, intermediate and minor ellipse axes)

4.

the CODE1 and MATL values of both the composite and the block must match (ie. only fresh composites from are used to interpolate a fresh block and vice versa for oxide)

The search ellipse orientations follow the kriging axes. The search ellipse dimensions allow the block being interpolated to ‘see’ two sections along strike and two holes up or down dip. The relative dimensions of the search ellipse axes approximate those of the ranges of the first structure.

To check the suitability of the search ellipses used, a search ellipse was created in MineSight for each run file allowing visual inspection of the composites used and kriging weights calculated for the block at the centre of the ellipse.

17.2.5.3 Bulk Density

There are no density data from within the model area. Consequently, bulk density values were assigned to the block model according to material type using values from data collected in the Central Area (see Table 17 above).

17.2.6 Mineral Resource Classification

Three solids were created enclosing regions of geological confidence (measured, indicated or inferred) and these three regions were in turn used to code the item GLCAT in the block model. The solids generally enclose areas of approximately equally spaced drilling, but also allow areas where there is reduced confidence in the geological interpretation to be reported to a lower confidence category. The Measured Resource solid is always surrounded by a halo of Indicated Resource. Similarly, the Indicated Resource solid is always surrounded by a halo of Inferred Resource.

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17.2.7 Results

Results for the Mineral Resource contained in the Northern model are provided in Table 13.

Oxide gold resources in the Hunt’s – Fosterville area are confined to remnant oxide mineralisation below and along strike from previously mined open pits. Primary gold resources are largely restricted to the upper 50 metres of the fresh zone.

The resources in the Hunt’s – Fosterville area are based on a substantial quantity of data, having been drilled to 25 by 25 metres in most places. The quality of the data used for resource estimation is variable with significant amounts of older RC samples from holes drilled with cross over hammers, spear sampled and suspected of significant down hole smearing. The poorer quality data is largely restricted to the oxide zone.

17.3 Southern Area

The southern model spans from the Harrier Pit area to Daley’s Hill pit, close to the southern margin of the Fosterville Mining Lease (MIN5404) as shown in Figure 9.

The Southern model was in existence before the Harrier Mine Model in the Central Area was operational. There is overlap between the two models between 6600mN and 7400mN and where this occurs, the Harrier Mine Model is solely used for reporting.

The Southern Model area has in the past been discussed with reference to three main northing subdivisions, shown below, rather than ore shoots or geological structures owing to insufficient understanding of the area:

1.

6100mN to 7400mN Harrier area

2.

5180mN to 6100mN Wirrawilla area

3.

 4500mN to 5180mN Daley’s Pit Area

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However, with the benefit of current diamond drilling, interpretation of the major mineralisation controlling structures, from the Harrier pit southwards to beneath the northern parts of the Daley’s Hill area, has been possible.

17.3.1 Area Geology

Within the Southern Model area, the controlling features include the Fosterville Fault and the footwall Harrier suite of faults which have variable reverse offsets. The Harrier basal fault, termed "Harrier Base", appears to have more displacement than the other faults and becomes more important to localising mineralisation southwards. The total displacement over the Harrier suite of faults is about 120 metres (Figure 20).

Figure 20: Schematic Wirrawilla geological cross section showing folded bedding, anticline and syncline axial planes, faulting and mineralisation

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Movement on the Fosterville Fault lessens from north (100 metres+) to south (~10 metres at Daley’s Hill) and becomes less important southwards with respect to mineralisation.

The east-west folding in the area, varies from gently southerly plunging in the north to moderate southerly plunging at Daley’s Hill in the south. Fold plunge is important as the mineralised west dipping fault geometry is controlled by eastern limbs of syncline fold plunges where the faults become unmineralised "bedded" laminated quartz features.

At Daley’s Hill, the Daley’s Hill Fault has an associated 10 metres of reverse fault movement and localises the bulk of gold mineralisation. In the Daley’s Hill pit, lesser well mineralized east-west structures occur in the eastern parts and several other poorly defined hangingwall mineralised fault structures are present in the western portions of the pit. Daley’s Hill is unusual in that late stage free primary gold is noted in two diamond holes.

The geology of the Southern Model area was reviewed by independent consultant Stephen King in 2004 (King, 2004) and the northern parts again in 2006 (King, 2006). Rod Boucher (consultant, Linex) has also contributed much to the stratigraphic-structural understanding of the area. The most recent geological interpretation is reported by Andrea Reed (2007).

17.3.2 Geological Models

Geological modelling undertaken was essentially identical to that described for the Northern Model above. Several iterations of resource modelling of the southern area were undertaken by Simon Hitchman (Northgate), previously reported in Hitchman (2006). A review of the 2006 resource work was undertaken by Scott Jackson from QG consultants (Jackson, 2007).

17.3.3 Domaining

Domaining of the Southern area was based on geological structure, orientation, material types and variography. The domains are:

1.

Fosterville Fault

2.

Harrier Splays

3.

Daley’s Hill N-S Faults

4.

Daley’s Hill E-W Faults

5.

Materials (Oxide, Transitional and Fresh)

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The mineralisation domains were created by firstly using a nominal 0.2 to 0.5 g/t Au outer limit for sectional strings in weathered areas and 0.5 to 1.0 g/t Au in unweathered mineralisation. These values used reflect natural breaks to the mineralisation.

The strings were then linked or extruded to form a three dimensional wireframe domain. The strings were generally extruded a maximum of half the drill spacing. This varied from as little as 5 metres, in well drilled pit locations, to 50 metres, where mineralisation extended over several 100 metre spaced drill sections. The mineralised wireframes are shown in Figure 21.

Figure 21: Longitudinal projection, looking west at Southern Model area. Shown are mineralisation wireframes, drill traces and open pits (yellow).

17.3.3.1 Fosterville Domain

Fosterville Fault related mineralisation occurs between 6350mN and 6550mN in the Johns Pit’s area. Fosterville Fault domain strikes N-S, and dips 70° west. The domain presently only persists across two 100 metre spaced drill sections in the Southern Area and is poorly defined.

17.3.3.2 Harrier Splays Domain

The Harrier Splays domain consist of seven individual fault strands that are interpreted to splay from the Fosterville Fault. The Harrier splays anastomose with one another, both along dip and along strike and have dips that tend to shallow westwards. The splays have average dips of 60° W and 350° N strike. The domain commences at the Harrier pit and gently plunges southwards to the Daley’s Pit area.

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17.3.3.3 Daley’s Hill Domains

The Daley’s Hill area has four separate northerly trending mineralised structures. At present these are treated identically in one domain owing to paucity of drilling information of the smaller mineralized envelopes. The domain has a variable Northerly strike, 70° W dips and 20° S plunge.

East-west mineralized structures occurring in the eastern parts of the pit are footwall to the main Daley’s Hill N-S structure. The Daley’s Hill E-W domain comprises 4 separate structures which trend 060° and dip 80°N.

17.3.3.4 Material Domains

Three ‘material’ domains were constructed, similar to that described previously in the Northern model. The domains are:

1.

Fresh (sulphide minerals completely un-oxidised)

2.

Transition (sulphide minerals may be partially oxidised, includes zones of mixed fresh and oxide)

3.

Oxide (sulphide minerals completely oxidised, Fe-carbonates largely oxidised)

The Transitional domain lower boundary is only an approximation because there is insufficient logging of the base of transition to allow a reasonable interpretation of this surface over the entire southern model. The base of transition was taken to be 5110mRL after comparison with drill data and results from open pit mining in the area.

Separate domains were constructed for transitional and fresh materials for metallurgical recovery studies. However, during block model interpolations, transitional and fresh material are treated as if they are the same material type, but coded appropriately in the block model for inventory purposes.

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17.3.4 Drilling Data

The drilling quality is variable in the Southern area and includes:

1.

RAB – Rotary Air Blast,

2.

Reverse circulation – Cross over hammer and face sampling hammer variants, and

3.

Diamond core – HQ and NQ2, often with RC pre-collars.

During drillhole data extraction for resource interpolations, the omission of RAB holes and one diamond hole was required owing to low quality sample techniques and incomplete assaying respectively. MineSight drill views were the primary tool used to identify data problems.

Subsequent to the drill data review process assay data were:

1.

Imported from the acQuire Exploration databases into MineSight using MineSight customisable parameter screens.

2.

Coded for mineralisation using 3-D gold wireframe solids.

Within the oxide open pit areas, the historical 5 metre blast holes are vertical and generally had one sample collected over a 5 metre length. These holes were used to aid interpretation, but were not used during subsequent kriging owing to sample quality and that the 5 metre sample lengths were in excess of the desired 2 metre composite lengths.

17.3.4.1 Compositing and Coding

Compositing and coding of drill holes was undertaken similar to the Central Area.

17.3.4.2 Variography

Owing to the coarseness of the exploration drill line spacing (nominal 100 metre) in the Southern Area, variography for the Fosterville Fault and Harrier Splays was adopted from the Central Model Areas for both the sulphide (fresh and transitional) and oxide gold domains. Similarly the lack of sulphur assays hampered variography studies.

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In the Daley’s Hill area where closer spaced drilling is available, variography work demonstrates relative nugget effect values of 50% and most of the variance in the first ~30m. The variogram models closely follow the expected geological controls with 20° southerly plunging shoots in 70° west dipping faults.

17.3.5 Resource Modelling

17.3.5.1 Block Models

The Southern Block Model was created to allow modelling of mineralisation south of 7,400mN. XYZ Block dimensions of 4 metres (east) by 10 metres (north) by 5 metres (RL) were used. This block size was chosen after consideration of the maximum drilling density (25 by 15 metres), mineralisation geometry (typical mineralisation width of 3m to 8m) and probable open pit mining methods.

17.3.5.2 Search Criteria

Gold grades were interpolated into blocks meeting the following criteria:

1.

greater than 1% of the block volume is inside one of the domain envelopes.

2.

blocks whose search ellipse includes at least 5 composites

3.

Blocks whose MATL code is set to Oxide (3), Transitional (2) or Fresh (1).

Similarly, only composites meeting the following criteria are used to interpolate any one block: 1. all composites (to a maximum of 30 composites) within the search ellipse dimensions and search area limits outlined in the table below. 2. where more than 30 composites lie within the search ellipse the 30 closest composites in ellipsoid space are used. 3. maximum of 6 composites are used from any split quadrant of the search ellipse.

1.

all composites (to a maximum of 30 composites) within the search ellipse dimensions and search area limits outlined in the table below.

2.

where more than 30 composites lie within the search ellipse the 30 closest composites in ellipsoid space are used.

3.

maximum of 6 composites are used from any split quadrant of the search ellipse.

4.

the CODE1 and MATL values of both the composite and the block must match (i.e. only fresh composites from are used to interpolate a fresh block and vice versa for oxide).

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The search ellipse orientations follow the kriging axes. The search ellipse dimensions allow the block being interpolated to ‘see’ two sections along strike and two holes up or down dip.

Sulphur grades were interpolated into blocks in the Harrier pit area prior to recommencement of sulphide gold mining in 2007. However, many historical RC holes did not have sulphur assays and reporting from the Harrier Model is preferable as it has many more grade control holes with sulphur assays.

17.3.5.3 Bulk Density

The bulk density profile established for the Central Area was taken as being appropriate for the Southern Model area given the similar rock types, levels of oxidation and identical mineralisation and gangue mineralogy (Table 17).

17.3.6 Mineral Resource Classification

Similar to the Central Area, three solids were created enclosing regions of geological confidence (Measured =1, Indicated=2 and Inferred=3) and these three regions were in turn used to code the item RSCAT in the block model. The solids generally enclose areas of approximately equally spaced drilling, but also allow areas where there is reduced confidence in the geological interpretation to be reported to a lower confidence category.

In areas of the Southern model away from the open pits, the diamond drilling is on nominal 100 metre north spaced drill sections with 50 metre down dip holes spacings, and for this drill density the mineralisation is broadly classified as Inferred. Beneath the open pits where the drill spacing is reduced to 10-20 metres north by 10-15 metres east, mineralisation is classified as Measured with a halo of Indicated.

The Daley’s Hill east-west structures are not well understood and as such this mineralisation is classified as Inferred.

17.3.7 Results

Results for the Mineral Resource contained in the Southern model are provided in Table 13.

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Small oxide gold resources exist in the John’s Pit and Daley’s Hill area and are confined to oxide mineralisation along strike from previously mined open pits in the top 40 metres. There is also a small pod of oxide to the south west of John’s pit that was not extracted during oxide mining.

The bulk of the sulphide resources within the Southern Model are based on 100 by 50 metre spaced diamond drilling supplemented by closer spaced, but lower quality face and cross over RC drilling. Infill drilling will be required to increase resource confidence from a largely Inferred mineral resource category.

17.4 Robbins Hill Area

The Robbins Hill Area is north-east of the Central Area and contains the O’Dwyer’s, Robbins Hill, Farley’s, Sharkey’s, Woolshed and Read’s oxide pits as shown in Figure 9. The controlling structural features in the area include a variety of north-trending west-dipping faults and failed anticline axes intruded by dykes. The Sharkey’s area has one of the few areas known on the lease where east dipping structures are significantly mineralised.

The geology of the area was assessed by Fosterville staff during diamond drilling activities between 2004 and 2007, reported by Chris Reed (2007) and reviewed twice by Stephen King (2005 and 2007). The area was also the subject of a study conducted by Chris Davis (Davis, 2006). Robbins Hill Model resource modelling conducted by Kerrin Allwood and Simon Hitchman is reported in Allwood (2006b) and Hitchman (2007). A further review of modelling in the Farley’s-Sharkey’s area is also reported in Allwood (2007).

17.4.1 Area Geology

17.4.1.1 O’Dwyer's to Robbins Hill Area – Geological Overview

The fault architecture of the O’Dwyer’s-Robbins Hill (ODW-RH) area is more complicated than that observed in the Fosterville Fault zone which generally has structural complexity on only one side of the Fosterville Fault.

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Three significant fold closures occur in the west of the area – the Robbins Hill Anticline and Syncline and the Trench Syncline (named after the trench it was exposed in). The latter is discussed further in the Farley’s-Sharkey’s section. The Robbins Hill Anticline and Syncline pair loses amplitude and wavelength southwards from a wavelength of around 100 metres in the north to become a small parasitic fold pair in the south of the Robbins Hill pit. The folds are complicated by fault structures and the anticline is intruded by a porphyry dyke (RH Porphyry) within the pit. The dyke intrudes along the axial plane and pinches out to the south. This is interpreted to coincide with the position where the fold pair has diminished and a consistently east dipping stratigraphy is present – the dyke is oblique to this stratigraphy and so cannot easily continue to the south.

There is considerable variation in fold plunge directions and amounts in the ODW-RH area. Mapped north plunging bedding-cleavage intersections are present in the north and west of the area whilst south plunging intersections occur in the rest of the area but the degree of south plunge can change across faults (the ODW South pit also has north plunging intersections which complicates the geometry further). The changes in fold plunge have a strong bearing on the complexity of the fault system as structures are commonly parallel to bedding. There is a change from grid north trending structures to more NNE trending structures moving northward which is probably intimately associated with fold morphology.

Two NNE trending structures (now beneath backfill) have been interpreted from Robbins Hill pit mapping at 12800mN. A zone of faulting links between the two zones and controls ore blocks which cross-cut bedding and may be interpreted as an extensional link structure in a component of sinistral shearing on the NNE faults.

The central and northern portions of the Robbin’s Hill pit have not yet been modelled.

In the ODW South and North pits the same west dipping fault structure is mineralised and has a curvilinear grid north trend. East of, and paralleling this fault, is an anticline structure which has mineralised porphyry dyke (ODW Porphyry) occupying the sub-vertical axial plane. The ODW porphyry occurs in the eastern portion of the ODW South pit and in the middle of the ODW Central pit. Several west dipping mineralised faults occur on both sides of the ODW Porphyry and outcrop in ODW Central and Eastern pits. The west dipping faults appear to stop at the ODW Porphyry (Figure 22).

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Figure 22: Plan view of ODW-RH area, showing oxide pits and mineralisation. In green are RH and ODW sub-vertical porphyry dykes. West dipping faults are in blue, browns, purples and magenta.

Northeast trending unconsolidated Murray Basin clays, sands and gravels mask the Ordovician basement in the northwest and southeast parts of the Robbin’s Hill Model area (see Figure 9).

17.4.1.2 Farley’s-Sharkey’s Area - Geological Overview

The Farley’s zone of mineralisation is hosted by an anticline-syncline pair. These folds represent a set of parasitic east verging structures on the eastern limb of the Trench Syncline (Figures 23, 24 and 25). The Trench syncline is a major north-south (350°) trending asymmetric syncline that plunges towards the south at about 15°. The eastern limb of the Trench syncline has a shallower dip than the western limb and as a consequence the parasitic folding and bedding on the eastern limb has an observed strike of approximately 340-350°. This asymmetry of the folding accounts for the atypical strike of the Farley’s Pit in plan view.

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Figure 23: Cross-section on 12300mN

On 12300mN, the Base Fault (red) is bedding parallel in the sedimentary black shale unit (blue) on the eastern limb of the Trench Syncline.

The Farley’s anticline and syncline parasitic folds are developed beneath a sedimentary black shale unit that wraps through the Trench syncline. This black shale unit can be seen in the Robbin’s Hill pit where it dips to the east. The stratigraphy above the black shale can be correlated between holes and appears to correlate with the stratigraphy above the black shale in the Robbin’s Hill pit.

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Figure 24: Cross-section on 12400mN

On 12400mN, parasitic folding (Anticline Hangingwall Splay and the Farley’s Syncline) is developed on the eastern limb of the Trench Syncline and is bounded by the Top Bounding Fault (Pink) at the base of the black shale unit (Blue) and the Base Fault (red). Sulphide mineralisation (gold lines adjacent to drill holes) is confined to the area beneath the black shale (and hydrocarbon alteration) where it wraps through the anticline and syncline hinges. A conceptual proposed pit (Nov 2006) outline is shown in dashed green.

The Top Bounding Fault is developed towards the base of black shale unit (Figure 24). This fault appears to have accommodated the movement in the black shale during folding and breaks out of the shale where it folds through the parasitic anticline and the Trench Syncline (Figures 23 and 24).

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On the eastern side, the parasitic folds are truncated by the Base Fault. This is major north-south trending fault zone that is broadly conformable to the footwall but is discordant to the hangingwall, i.e. it separates the asymmetry of the western limb of the Trench syncline from north-south trending west dipping bedding in the footwall. Given the form of this fault, it is interpreted as a major dislocation surface that has developed beneath the parasitic folding. The parasitic folding is therefore confined to the space between the Base Fault and the Top Bounding Fault in the black shale unit.

Figure 25: Cross-section on 12650mN

On 12650mN, the main sulphide zone associated with the faulted Farley’s Anticline steps up onto the faulted Trench Syncline. This corresponds with the Trench Syncline hinge being faulted out on the Top Bounding Fault. The parasitic folding observed to the south has now become more open and the base fault has steepened markedly with the Farley’s Syncline hinge now seen in the both the Footwall and Hangingwall of the fault. The kink zone in the Base Fault (and therefore the position of the Farley’s Footwall Syncline) controls the position of the high grade zone intersected on the base fault.

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The zone between the Base Fault and Top Bounding Faults represents a zone of increased deformation with shearing and stockworking developed between the two faults but focussed on the two parasitic fold hinges. Within this zone, deformation has occurred predominately as puggy faulting with little actual displacement on any one fault. In the sandstone units, stockworking is well developed hangingwall and footwall to the puggy faults. It is this puggy faulting and quartz stockworking that hosts the bulk of the sulphide mineralisation. Minor higher grade mineralisation occurs on the Base Fault where it truncates the anticline hinge.

Moving to the north, the folding becomes more open and the two parasitic folds adopt a more regional north-south trend. This transition occurs at about 12650mN and corresponds with a plunge reversal in the Trench Syncline. In the area of this reversal, mineralisation steps from the anticline-syncline pair to the faulted Trench Syncline (Figure 25). This transition has been termed the culmination transfer zone.

17.4.2 Geological Models

Geological modelling undertaken was essentially identical to that described for the Northern and Southern Models described above.

17.4.3 Domaining

Basic statistics and variographic analysis was completed on the interpreted mineralisation wireframes in the O’Dwyer’s-Robbin’s Hill area and also in the Farley’s-Sharkey’s area. Each area was subdivided into two domains chosen based on (in order of decreasing importance) geology, variography and statistics.

In the O’Dwyer’s-Robbin’s Hill area, oxide and sulphide mineralisation was grouped into single domains for the Porphyry and the Faults domains because there is very little difference in the statistics of the oxide and sulphide mineralisation for these domains.

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In the Farley’s-Sharkey’s area there is sufficient statistical differences between the oxide and combined transitional-fresh zones to isolate these

The domains include:

1.

– ODW Porphyries

2.

RH-ODW Faults

3.

Farley’s Faults

4.

Sharkey’s East Dipper

5.

Materials (Oxide, Transitional and Fresh)

17.4.3.1 Gold Domains

The RH-ODW Porphyry domain encompasses all porphyry hosted mineralisation. The Porphyry domain is low grade with rare extreme gold grades occurring on the contact. The RH-ODW Faults domain includes the O’Dwyer’s Fault, O’Dwyer’s East A, B, and C, and the Robbin’s Hill East A, B and C sub-domains. All of these sub-domains strike approximately north – south and dip 50° to 70° west. The Faults domain is characterised by low grade, narrow but continuous mineralisation.

The Farley’s Faults domain dip 50°-70° west and show a change in strike from 345° at Farley’s pit to 360° south of Sharkey’s pit. The faults are characterised by low grades with rare high grade values.

The Sharkey’s East Dipper domain dips 45° east, strikes north and generally has low continuous grades in the fresh domain, but is more erratic in the oxide materials.

17.4.3.2 Oxidation Domains

Four ‘material’ domains were constructed, similar to that described for the Northern and Southern Models, although only gold mineralisation was interpolated in three of these.

The four domains are:

1.

Fresh (sulphide minerals completely un-oxidised)

2.

Transition (sulphide minerals may be partially oxidixed, includes zones of mixed fresh and oxide)

3.

Oxide (sulphide minerals completely oxidixed, Fe-carbonates largely oxidised)

4.

Alluvium (near surface transported material, generally barren of gold, largely clay, free digging)

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17.4.4 Drilling Data

The quality of the drilling is variable in the Robbin’s Hill area with drilling spans the years 1989 to 2007. Up until 2001, drilling was focussed on oxide heap leach targets and as such cheaper less precise drilling methods were used and dominate the dataset. After 2004, diamond holes were used to aid structural interpretation and often RC pre-collars were diamond tailed.

The model uses more than a total of 1600 holes of which about 95% are RC holes and 5% are NQ2 and HQ diamond core holes. Drill data omitted where there was uncertainty of coordinates, dubious down hole surveys and grade or geological mismatch. MineSight drill views were the primary tool used to identify data grade and geological mismatches.

Subsequent to the drill data review process assay data were:

1.

Imported from the acQuire Exploration databases into MineSight using MineSight customisable parameter screens.

2.

Coded for mineralisation using 3-D gold wireframe solids.

17.4.4.1 Compositing and Coding

Similar to the Central Area, coded Robbin’s Hill Model area drill data was composited to 2 metre lengths starting from the point at which the drillhole enters the mineralisation envelope. If the final composite was less than 1.0 metres it was added to the previous composite making a composite with length between 2.0 and 3.0 metres. Final composites between 1.0 and 2.0 metres in length were left as is.

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Two metre composites were chosen to reflect the anticipated minimum mining width, to allow across strike variability to be maintained within the data and because the vast majority of RC drilling samples is two metres long.

17.4.4.2 Variography

In all domains, the nugget effect (46 to 59%) is typical of gold deposits, but higher than other mineralisation domains in the Fosterville Goldfield. Typically low nugget effects elsewhere at Fosterville reflect the fine grained, disseminated nature of the sulphide minerals hosting the elements analysed and are confirmed by the very low variability exhibited in assay QAQC data. The higher nugget effects modelled for these domains may reflect some mixing of populations, possibly owing to re-mobilisation of gold by weathering resulting in erratically distributed extreme gold grades.

The longer range structures in the RH-ODW area possibly reflect high grade zones occurring where faults intersect the quartz porphyry dykes. The variogram models closely follow the expected geological controls with flat to shallowly south plunging shoots in steeply west dipping faults and sub vertical porphyry contact zones.

17.4.5 Resource Modelling

17.4.5.1 Block Models

The Robbins Hill block model was created in 2005 and has sufficient extents to contain all drilled mineralisation beneath the open pits in the area. Previously, several smaller block models were used to inventory mineralisation for the oxide pits in the area. These models had differing block dimensions and orientations from one another and so combining them into a single unified model was not possible.

The Robbins Hill Model has block dimensions of 4 metres (EW) by 10 metres (NS) by 5 metres (RL). The four metre width was chosen as it is approximates the minimum mining width for both open pit and underground mining. The ten metre north – south block dimension is half the section spacing in the most densely drilled areas. The five metre vertical block dimension is the likely mining bench height and allows sufficient resolution for future pit optimisation. Block dimensions are identical to those of the Southern and Northern Models.

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Resource modelling in the O’Dwyer’s and southern Robbin’s Hill area was undertaken by Kerrin Allwood (2006b) although the central and northern portions of the Robbin’s Hill pit mineralisation still remain un-modelled owing to priorities and are not reported. Similarly the gold mineralisation for the Woolshed and Read’s pits has not been remodelled but previous December 2001 modelling results are available and were used in compiling resource reports for the Robbins Hill Model area. Resource modelling of the Farley’s to Sharkey’s area was undertaken by Simon Hitchman between 2005 and 2007 and reviewed by Kerrin Allwood (2007).

17.4.5.2 Search Criteria

Gold and sulphur grades are only interpolated into blocks meeting the following criteria:

1.

Greater than 1% of the block volume is inside one of the domain envelopes.

2.

Blocks whose search ellipse includes at least 5 composites.

3.

Blocks whose MATL code is set to Oxide (3), Transitional (2) or Fresh (1).

Similarly, only composites meeting the following criteria are used to interpolate any one block:

1.

All composites (to a maximum of 15 composites in the O’Dwyer’s-Robbin’s Hill area and 30 in the Farley’s-Sharkey’s area) within the search ellipse dimensions and search area limits outlined in the table below.

2.

Where more than 15/ 30 composites lie within the search ellipse the 30 closest composites in ellipsoid space are used.

3.

Maximum of 6 composites are used from any split-quadrant of the search ellipse.

4.

the CODE1 and MATL values of both the composite and the block must match (ie. only composites from within the same mineralisation envelope and the same oxidation material are used to interpolate a block). However, no distinction is made between Transitional and Fresh materials and both are treated as if they are fresh.

5.

1-3m in length

6.

Within the specified search ellipse (dimensions and orientation defined by block CODE1 value)

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The search ellipse orientations follow the kriging axes. The search ellipse dimensions allow the block being interpolated to ‘see’ two sections along strike and two holes up or down dip. To check the suitability of the search ellipses used, search ellipses were created in MineSight to allow visual inspection of the composites to be used.

17.4.5.3 Bulk Density

The bulk density profile established for the Central Area was taken as being appropriate for the Robbins Hill Model area given the similar rock types, levels of oxidation and identical mineralisation and gangue mineralogy (Table 17).

17.4.6 Mineral Resource Classification

No resources in the Robbins Hill Area have been categorised as Measured owing to uncertainties in the quality of the largely historical data used to construct this model.

Two solids were created enclosing regions of geological confidence (Indicated or Inferred) and these regions were in turn used to code the item RSCAT in the block model. The solids generally enclose areas of approximately equally spaced drilling, but also allow areas where there is reduced confidence in the geological interpretation to be reported to a lower confidence category. The Indicated Resource solid is always surrounded by a halo of Inferred Resource.

17.4.7 Results

Consolidated results for the Robbins Hill Model are provided in Table 13.

Oxide gold resources exist in the Robbins Hill Model area, notably east of Sharkey’s pit where exploration drilling in 2007 discovered shallow oxide mineralisation. Elsewhere remnant low grade oxide gold mineralisation below and along strike from previously mined open pits.

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Resources in the Farley’s-Sharkey’s area are based modern face sampling RC methods and substantial diamond drilling and as such the geological information is better than elsewhere in the modelled area.

18. Other Relevant Data and Information

Not applicable.

19. Interpretation and Conclusions

The authors have made the following interpretations and conclusions:

1.

The understanding of the geological controls on mineralisation at Fosterville is high. Primary mineralisation is structurally controlled with high grade zones localised by the geometric relationship between bedding and faulting. This predictive model has lead to considerable exploration success in following the down-plunge extensions of high grade mineralisation.

2.

By the same token, this understanding has lead to the development of robust geological and resource models underpinning the Mineral Resource and Mineral Reserve estimates. The relationship between mineralisation and the controlling structural/stratigraphic architecture means that quality geological interpretation is critical to producing quality resource/reserve estimates.

3.

The modifying factors used to convert the Mineral Resources to Mineral Reserves have been refined with the operating experience gained since underground production commenced in September 2006. In particular, the robustness of the mining recovery and dilution estimates have improved with experience relative to the pre-mining assessments.

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20. Recommendations

The following recommendations are made:

1.

Further mine lease exploration/resource development activities should be pursued. Given the strong understanding of geological controls on mineralisation, this would be considered likely to yield additional resources and reserves. Particular areas that are recommended to focus upon are the Wirrawilla (5300] 6100mN) and Daley’s Hill (4700] 5200mN) zones in the Southern Area and the Robbin’s Hill North area (12500] 12900mN) in the northern portion of the Fosterville Mine Lease. . Proposed and costed RC and diamond drilling programs for each of these areas are shown in Table 20 below.

Target / Drill Program	RC	DDH	Estimated Cost
	(Metres)	(Metres)	(AUD Millions)
Wirrawilla	4,900	12,300	2.9
Daley’s Hill	0	1,200	0.3
Robbin's Hill North	0	2,700	0.6

Table 20: Proposed drilling programs for 2008

2.

The infill drilling program should be continued with an aim to maintain at least 12 months of reserves drilled out to 25 metre centres (or closer where necessary). Given the south plunging geometry of the Phoenix orebody, this is being conducted from a dedicated development drive in the hangingwall of the Fosterville Fault (western side). The current infill drilling budget for 2008 includes 30,000 metres of drilling at an estimated cost of approximately A$3.6 million. As the decline and mining front continues to move south and lower, further hangingwall drives will be required to be developed. This work and the associated drilling have not yet been costed in detail.

3.

With this additional drilling data and further ongoing operational experience, it is recommended that mining recovery and dilution factors are reviewed and refined on an ongoing basis.

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21. References

Allwood, K, 2003. Fosterville Gold Project April 2003 Resource Estimate. Unpublished internal report.

Allwood, K, 2005. May 2005 Resource Model Description. Unpublished internal report.

Allwood, K, 2006a. Perseverance Exploration Limited Fosterville Gold Project June 30 2006 Fosterville ML Resource Estimates section 1 - central area. Unpublished report by Geomodelling.

Allwood, K, 2006b. June 30 2006 Fosterville ML Resource Model Descriptions Section 3 – O’Dwyer’s Area. Unpublished report by Geomodelling.

Allwood, K, 2007. Review of Resource Modelling Procedures for the Farleys Area, Unpublished report by Geomodelling.

Allwood, K, 2008. Hunt’s – Fosterville Resource Estimate (Northern Block Model), Fosterville Gold Project. Unpublished report by Geomodelling.

Bureau of Meteorology, 2008. http://www.bom.gov.au/, March.

Crase, N, 2002. Brief notes on the review of Quality Assurance and Quality Control insampling and assaying at Fosterville, May 2002. Unpublished memorandum by SMP Consultants.

Davis, C, 2006. Structures and Alteration of the Robbin’s Hill Pit Area. Unpublished University of Melbourne Batchelor of Science Honours Thesis.

Hitchman, S, 2006. June 30 2006 Fosterville ML Resource Model Descriptions Section 2 - Southern Areas, July, 2006. Unpublished internal report.

Hitchman, S, 2007. Farley’s Area Resource Work - May 2007. Unpublished internal report.

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Jackson, S, 2007. Review of Fosterville and Golden Gift 3 & 9 Resource Estimates. Unpublished report by QG Consulting.

Kelemen, T, 2004. Perseverance – Data Systems Review, March 2004. Unpublished report  by IO Digital Systems.

King, S, 2004. Geology of the Harrington’s Hill to Daley’s Hill Area with Special Reference to Daley’s Hill Pit Geology. Unpublished report by Solid Geology.

King, S, 2005. Structural Compilation of Drilling and Geological Mapping in the Robbin’s Hill – Sharkey's Area, Fosterville Gold Project, Victoria. Unpublished report by Solid Geology.

King, S, 2006. Structural Interpretation of Cross-sections, Between 6700N and 5900N (Harrier Splay) Fosterville Gold Project, Victoria. Unpublished report by Solid Geology.

King, S, 2007a. Structural Interpretation of The Fosterville and Hunt’s Pit Area 9900N – 11400N Fosterville Gold Project, Victoria. Unpublished report by Solid Geology.

King, S, 2007b. Structural Interpretation of Fault System around Farleys Pit 12300N – 12550N Fosterville Gold Project, Victoria. Unpublished report by Solid Geology.

McConville, F, 2006. Perseverance - Data Systems Review, September 2006. Unpublished report by IO Digital Systems.

Perseverance, 1997. Sulphide Open Pit Feasibility Study. Unpublished internal report.

Perseverance, 2000. Sulphide Open Pit Feasibility Study. Unpublished internal report.

Perseverance, 2003. Fosterville Bankable Feasibility Study. Unpublished internal report.

Reed, C, 2007. Farley’s Deposit Geological Summary, April 2007. Unpublished internal report.

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Roberts C, Jackson T, Allwood K, Shawcross M, Story J, Barbetti L, Tielen R, Boucher R and Norris N, 2003. Fosterville - Rise of the Phoenix, the emerging goldfield at Fosterville, in NewGenGold 2003 Conference Proceedings, (Louthean Media: Perth).

Stewart, M, 2007. Notes on QAQC processes and On-Site Laboratory visit. Unpublished report by QG Consulting.

Vandenberg A H M, Willman, C E, Maher, S, Simons, B A, Cayley, R A, Taylor, D H, Morand, V J, Moore, D H, & Radojkovic, A, 2000. The Tasman Fold Belt in Victoria, in Geological Survey of Victoria Special Publication.

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

22.1 Mining Operations

Since the completion of the Harrier Open Cut mine in early December 2007, the sole source of ore has been the underground operations. Although potential exists, there are no advanced plans for further open cut mining at Fosterville. This is reflected in the current Life of Mine (LOM) plan which is based solely upon mining underground ore.

The underground mine commenced declining in March 2006 and with production first recorded in September 2006. Development and stoping have commenced in the Falcon and Phoenix orebodies with development yet to commence in the Ellesmere and Kink orebodies. The Phoenix orebody is being mined using open stoping methods – longitudinal retreat in areas <10 metres in width and transverse where the width exceeds 10 metres. The Falcon orebody is being mined by a modified longitudinal sublevel caving method with oxide from surface being drawn down to support and confine the hangingwall. The Ellesmere and Kink orebodies are currently planned to be mined by longitudinal retreat open stoping aside from the section of Ellesmere immediately under the open cut which is planned to be caved (see Figure 18 and Table 18 above). In all cases, the planned levels are 20 metres apart vertical (see Figure 26).

Figure 26: Longitudinal section showing actual and proposed mining layout as at November 2007.

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Mining is conducted using a conventional fleet including jumbos, production drills, loaders, trucks and ancillary equipment. Current mining is undertaken by a mining contractor (MG Mining) which has a whole-of-mine contract. In March 2008, it was announced that Fosterville will be moving to owner mining with the transition expected to take several months.

The processing path for the ore involves crushing and grinding followed by flotation, bacterial oxidation and CIL circuits. The bacterial oxidation process uses BIOX technology, operated under licence from Goldfields. The flowsheet can be seen in Figure 27. The processing capacity is approximately 1 Mtpa.

Figure 27: Schematic Ore Treatment Flowsheet

The production forecast contained in Table 21 was finalised in August 2007 as part of the latest Life-of-Mine (LOM) model. This model will also be discussed in sections 22.7 and 22.8 below. This LOM model was for the period July 1, 2007 to June 30, 2014 and used reserves only. Note that actual production for the period from July 1 to December 31 2007 was 360,280t at 3.72 g/t Au.

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				Year ended 30 June
								Total
	2008	2009	2010	2011	2012	2013	2014	2008-2014
Underground Production
Tonnes mined (t)	581,771	727,744	956,166	1,160,882	858,475	607,850	704,020	5,596,908
Head grade (g/t Au)	4.75	4.84	4.39	4.35	4.18	4.55	4.22	4.44
Surface Production
Tonnes mined (t)	118,923							118,923
Head grade (g/t Au)	2.76							2.76
Total Production
Tonnes mined (t)	700,694	727,744	956,166	1,160,882	858,475	607,850	704,020	5,715,831
Head grade (g/t Au)	4.41	4.84	4.39	4.35	4.18	4.55	4.22	4.41

Table 21: Production forecast for the years 2008-2014 (LOM plan).

22.2 Recoverability

Section 16 in this report contains a description of the metallurgical testwork that has been conducted for the reserves at Fosterville.

22.3 Markets

Fosterville produces gold doré bars mine site, which are transported to AGR Matthey in Western Australia and refined to produce gold bullion. The gold bullion is sold over the counter according to the corporate treasury policy through either AGR Matthey or an Australian based bank.

22.4 Contracts

Fosterville Gold Mine is subject to a licence fee following a licence agreement entered into with Minsaco by Perseverance in 2003. Minsaco has a licence from the proprietor to implement a process known as the BIOX^® process in Australia whereby micro-organisms are used in the oxidation of certain gold bearing sulphidic minerals in order to facilitate gold recovery. Perseverance has agreed to pay a licence fee to Minsaco calculated as an amount determined by multiplying the number of ounces of gold produced from FGM treated through the BIOX^® Plant by $1.40. The licence fee was payable from the date of commencement of operations at FGM and shall terminate when 1,500,000 ounces of gold in the aggregate has been produced from FGM treated at the BIOX^® plant.

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The authors are not aware of any agreements that are not within normal market parameters.

22.5 Environmental Considerations

The environmental bond is currently set at A$5,131,523 and is reviewed annually with the Department of Primary Industries in Victoria. Rehabilitation is undertaken progressively at the Fosterville Gold Mine but the environmental bond is only reduced on establishment of the rehabilitation which is not considered to have occurred until at least 5 years after rehabilitation has occurred.

22.6 Taxes

Fosterville is currently subject to the following taxes and duties:

• Victorian Payroll tax of payable at the rate of 5.05% calculated on wages paid by Fosterville to its employees. Fosterville is liable for Victorian Payroll Tax when its total Australian wages exceeds the Victorian general deduction threshold of A$45,833 a month or A$550,000 over a full financial year; and / or when grouped with other businesses of the Corporation, the combined Australian wages of the group exceed the general deduction threshold level.

• Minor taxes and duties including land tax, insurance duty, mortgages duty, motor vehicle duty debits tax and local council rates.

The Australian parent entity, Northgate Australian Ventures Corporation Pty Ltd is currently subject to the following taxes:

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• Federal Income tax is levied on the taxable income of the Corporation at a rate of 30%. In general terms, taxable income is calculated on assessable income less any allowable deductions.

• Capital Gains Tax (CGT) is paid on any capital gain that the Corporation includes in its annual income tax return. CGT is not a separate tax; it’s a component of income tax.

• Fringe benefits tax (FBT) is payable for benefits paid to an employee or the employee’s associate by the Corporation. FBT is separate from income tax and is based on the taxable value of the various benefits provided.

• Goods and Services Tax (GST) is a broad-based tax of 10 per cent on the sale of most goods and services and other things in Australia. Being registered for GST enables the Corporation to claim input tax credits for the GST included in the purchase price of goods and services used in the business.

• Victorian Payroll tax of payable at the rate of 5.05% calculated on wages paid it and its subsidiaries to its employees. Corporation is liable for Victorian Payroll Tax when its total Australian wages exceeds the Victorian general deduction threshold of A$45,833 a month or A$550,000 over a full financial year; and / or when grouped with other businesses, the combined Australian wages of the group exceed the general deduction threshold level.

• Minor taxes and duties including land tax, insurance duty, mortgages duty, motor vehicle duty debits tax and local council rates.

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22.7 Capital and Operating Estimates

EXECUTIVE SUMMARY	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9	TOTAL
	2008	2009	2010	2011	2012	2013	2014	2008-2014
Operating Costs (Cash Costs)
Mining - Underground	27,845	34,145	41,647	44,154	24,631	17,398	22,328	212,149
Deferred Mining - Underground	(2,150)	-	-	-	-	-	-	(2,150)
Mining - Surface	809	-	-	-	-	-	-	809
Deferred Mining - Surf. & Inventory adjustment	591	-	-	-	-	-	6,795	7,386
Mining - Mine Geology	4,546	4,672	4,540	3,321	2,400	2,186	1,955	23,619
Processing	14,253	13,134	17,146	17,944	13,870	8,457	8,692	93,495
Maintenance	4,432	3,330	5,184	5,074	5,074	4,642	4,338	32,074
Environment	260	247	245	245	252	228	204	1,681
OH&S	619	620	622	622	622	610	567	4,284
HR	231	273	289	289	289	231	76	1,678
Community	82	80	80	80	80	64	48	513
Administration	2,961	3,060	3,075	3,075	3,075	2,896	2,592	20,732
Total operating cash costs	54,479	59,560	72,827	74,804	50,292	36,712	47,595	396,270
Deferred mining & Inventory Adj Incl Above	(591)	-	-	-	-	-	(7,295)	(7,886)
Operating Costs (Cash flow)	53,888	59,560	72,827	74,804	50,292	36,712	40,301	388,384
Capital Costs
Mining	16,510	18,541	10,241	3,089	2,280	-	-	50,661
Infrastructure	1,980	3,300	2,900	1,400	1,200	200	1,200	12,180
Plant & Equipment	1,246	411	80	80	80	20	-	1,917
Other	-	-	-	-	-	-	-	-
Pre-Operations Operating Costs	2,150	-	-	-	-	-	-	2,150
Land & Buildings	107	130	-	-	-	-	-	237
OH&S	35	175	245	400	400	350	1,200	2,805
Processing Plant	2,317	6,305	3,800	-	-	-	(20,000)	(7,578)
Total Capital	24,345	28,863	17,266	4,969	3,960	570	(17,600)	62,373
Total Operating and Capital Costs (Cash Flow)	78,233	88,423	90,093	79,773	54,252	37,282	22,701	450,757

Table 22: Capital and operating cost estimates from the LOM plan completed in August 2007.

22.8 Economic Analysis

Table 23 sets forth an economic analysis for the FGM reflecting a cash flow forecast on an annual basis from 1 July 2007. In addition, certain sensitivity analyses are described in Figure 28 below. The cash flow forecast in Table 23 has been based upon only the mineral reserve estimate for the FGM as reflected in this report, representing approximately a seven year life. Gold price assumption is A$750/oz which is consistent with the price used to estimate the reserves.

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EXECUTIVE SUMMARY	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9	TOTAL
	2008	2009	2010	2011	2012	2013	2014	2008-2014
Production
Underground Production
Tonnes mined	581,771	727,744	956,166	1,160,882	858,475	607,850	704,020	5,596,908
Head grade	4.75	4.84	4.39	4.35	4.18	4.55	4.22	4.44
Production (oz)	88,766	113,319	134,983	162,297	115,490	88,885	95,530	799,269
Surface Production
Tonnes mined	118,923	-	-	-	-	-	-	118,923
Head grade	2.76	-	-	-	-	-	-	2.76
Production (oz)	10,560	-	-	-	-	-	-	10,560
Total Production
Tonnes mined	700,694	727,744	956,166	1,160,882	858,475	607,850	704,020	5,715,831
Head grade	4.41	4.84	4.39	4.35	4.18	4.55	4.22	4.41
Production (oz)	99,326	113,319	134,983	162,297	115,490	88,885	95,530	809,829
Surface ore ratio (%)	17%	0%	0%	0%	0%	0%	0%	2%
Total Processed
Tonnes processed	815,692	733,012	956,284	1,011,036	966,762	649,613	791,804	5,924,205
Head grade	4.32	4.84	4.39	4.29	4.26	4.43	4.21	4.25
Production (oz)	113,201	114,075	135,050	139,456	132,428	92,574	107,164	833,948
Recovery	79%	85%	87%	90%	90%	90%	90%	87%
Recovered gold (oz)	89,787	96,964	117,494	125,510	119,185	83,317	96,448	728,704
Revenue	$000	$000	$000	$000	$000	$000	$000	$000
Gold Sales	67,340	72,723	88,120	94,133	89,389	62,488	72,336	546,528
Interest Received and other revenue	-	-	6,887	3,962	-	-	2,615	13,464
Total	67,340	72,723	95,007	98,095	89,389	62,488	74,951	559,992
Spot Price (A$)	$750	$750	$750	$750	$750	$750	$750	$750
Operating Costs (Cash Costs)
Mining - Underground	27,845	34,145	41,647	44,154	24,631	17,398	22,328	212,149
Deferred Mining - Underground	(2,150)	-	-	-	-	-	-	(2,150)
Mining - Surface	809	-	-	-	-	-	-	809
Deferred Mining - Surf. & Inventory adjustment	591	-	-	-	-	-	6,795	7,386
Mining - Mine Geology	4,546	4,672	4,540	3,321	2,400	2,186	1,955	23,619
Processing	14,253	13,134	17,146	17,944	13,870	8,457	8,692	93,495
Maintenance	4,432	3,330	5,184	5,074	5,074	4,642	4,338	32,074
Environment	260	247	245	245	252	228	204	1,681
OH&S	619	620	622	622	622	610	567	4,284
HR	231	273	289	289	289	231	76	1,678
Community	82	80	80	80	80	64	48	513
Administration	2,961	3,060	3,075	3,075	3,075	2,896	2,592	20,732
Total operating cash costs	54,479	59,560	72,827	74,804	50,292	36,712	47,595	396,270
Deferred mining & Inventory Adj Incl Above	(591)	-	-	-	-	-	(7,295)	(7,886)
Operating Costs (Cash flow)	53,888	59,560	72,827	74,804	50,292	36,712	40,301	388,384
Operating Cash Flow	13,452	13,162	22,180	23,291	39,096	25,776	34,650	171,608
Capital Costs
Mining	16,510	18,541	10,241	3,089	2,280	-	-	50,661
Infrastructure	1,980	3,300	2,900	1,400	1,200	200	1,200	12,180
Plant & Equipment	1,246	411	80	80	80	20	-	1,917
Other	-	-	-	-	-	-	-	-
Pre-Operations Operating Costs	2,150	-	-	-	-	-	-	2,150
Land & Buildings	107	130	-	-	-	-	-	237
OH&S	35	175	245	400	400	350	1,200	2,805
Processing Plant	2,317	6,305	3,800	-	-	-	(20,000)	(7,578)
Total Capital	24,345	28,863	17,266	4,969	3,960	570	(17,600)	62,373
Total Operating and Capital Costs (Cash Flow)	78,233	88,423	90,093	79,773	54,252	37,282	22,701	450,757
Operating and Capital Cash Margin	(10,893)	(15,701)	4,914	18,322	35,136	25,206	52,250	109,235
Resource Definition Costs	1,200	-	-	-	-	-	-	1,200
Borrowing Costs	167	46	39	28	17	6	-	303
Cash Margin Before Tax	(12,260)	(15,747)	4,875	18,294	35,119	25,200	52,250	107,732

Table 23: Cash flow summary for Fosterville for the period 2008-2014.

Figure 28 below shows a sensitivity analysis with variations in price, costs, grade and recovery.

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Figure 28: Sensitivity analysis.

22.9 Payback

There is no outstanding debt on the property.

22.10 Mine Life

As contained in the production forecast in section 22.1 (Table 21), the current Fosterville mine life continues until 2014 based upon current Mineral Reserves.

In the opinion of the authors, the exploration potential on the mine lease is high. Given the strong understanding of geological controls on mineralisation, exploration/resource development activities would be considered likely to yield additional resources and reserves. Particular areas that are recommended to focus upon are the Wirrawilla zone (5300-6100mN) in the Southern Area and the down-plunge extension of the Phoenix orebody beyond the current southern terminus of Mineral Reserves at approximately 7400mN. Detailed programs including costing have not yet been prepared.

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23. Date and Signature Page

This Technical Report is effective as of March 25, 2008.

118

24. Certificates of Qualified Persons

119

Certificate of Qualified Person

Name: Simon Hitchman

Address: c/o Fosterville Gold Mine, McCormicks Road, Fosterville, Victoria, Australia, 3557

I, Simon Hitchman, MAusIMM, MAIG do hereby certify that:

1.

I am employed as the Exploration Manager (Acting) at the Fosterville Gold Mine.

2.

I have been jointly responsible for the preparation of the report titled "Technical Report on Fosterville Gold Mine", dated March 25, 2008, other than with respect to the Mineral Resource estimate for the Northern, Southern and Robbins Hill Areas for which I am wholly responsible as the "qualified person".

3.

I graduated from Melbourne University with a Bachelor of Science with Honours degree in 1986.

4.

I am a member in good standing of the Australasian Institute of Mining and Metallurgy and the Australian Institute of Geoscientists and have over 20 years of professional experience.

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 fulfil the requirements to be a "qualified person" for the purposes of NI 43-101.

6.

 I have worked continuously at the Fosterville Gold Mine since July 2004.

7.

I am not independent of the issuer applying all of the tests in section 1.4 of National Instrument 43-101.

8.

I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

9.

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

10.

I consent to the filing of the Technical Report with any stock exchange and any other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their website accessible by the public, of the Technical Report.

Dated 5^th June, 2008

120

Certificate of Qualified Person

Name: Ian Holland

Address: c/o Fosterville Gold Mine, McCormicks Road, Fosterville, Victoria, Australia, 3557

I, Ian Holland, MAusIMM do hereby certify that:

1.

 I am employed as the Principal Mine Geologist at Fosterville Gold Mine.

2.

 I have been jointly responsible for the preparation of the report titled "Technical Report on Fosterville Gold Mine", dated March 25, 2008, other than with respect to the Mineral Resource estimate for the Central Area for which I am wholly responsible as the "qualified person".

3.

 I graduated from James Cook University with a Bachelor of Science degree in 1996 and a Master of Minerals Geoscience degree in 2007.

4.

I am a member in good standing of the Australasian Institute of Mining and Metallurgy and have over 12 years of professional experience.

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 fulfil the requirements to be a "qualified person" for the purposes of NI 43-101.

6.

 I have worked continuously at the Fosterville Gold Mine since July 2007.

7.

 I am not independent of the issuer applying all of the tests in section 1.4 of National Instrument 43-101.

8.

I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

9.

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

10.

I consent to the filing of the Technical Report with any stock exchange and any other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their website accessible by the public, of the Technical Report.

Dated 5^th June, 2008

121

Certificate of Qualified Person

Name: Brad Evans

Address: c/o Fosterville Gold Mine, McCormicks Road, Fosterville, Victoria, Australia, 3557

I, Brad Evans, MAusIMM do hereby certify that:

1.

 I am employed by Mining Plus Pty Ltd as a consulting mining engineer since January 2008. I was previously employed as the Senior Mining Engineer for the Fosterville Gold Mine from September 2006 to January 2008.

2.

I have been jointly responsible for the preparation of the report titled "Technical Report on Fosterville Gold Mine", dated March 25, 2008, other than with respect to the Mineral Reserve estimate for which I am wholly responsible as the "qualified person".

3.

 I graduated from the University of Ballarat with a Bachelor of Engineering degree in 1998.

4.

I am a member in good standing of the Australasian Institute of Mining and Metallurgy and have over 10 years of professional experience.

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 fulfil the requirements to be a "qualified person" for the purposes of NI 43-101.

6.

I have worked continuously at the Fosterville Gold Mine since September 2006.

7.

 I am not independent of the issuer applying all of the tests in section 1.4 of National Instrument 43-101.

8.

I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

9.

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

10.

I consent to the filing of the Technical Report with any stock exchange and any other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their website accessible by the public, of the Technical Report.

Dated 5^th June, 2008

122