ex99-1.htm

TABLE OF CONTENTS

1.0	Summary	1
1.1	Executive Summary	1
1.1.1	Recommendations	2
1.2	Introduction	2
1.3	Property Exploration	4
1.4	Geological Setting	4
1.5	Pinson Project – Geology, Alteration, and Mineralization	5
1.6	Mineral Resource Estimation	5
1.7	Conclusions	7
1.8	Recommendations	8
1.9	Program and Budget	9
2.0	Introduction and Terms of Reference	10
2.1	Introduction	10
2.2.	Terms of Reference	10
3.0	Disclaimer	12
4.0	Property Description and Location	13
4.1	Location	13
4.2	Land Status and Agreements	14
4.2.1	Pinson Mining Company Agreement	15
4.2.2	Underlying Agreements – Unpatented Federal Mining Claims	18
4.2.3	Underlying Agreements – Patented Fee Lands	19
4.2.4	Underlying Agreements – Other Agreements	19
4.3	Royalty Summary	21
5.0	Accessibility, Climate, Local Resources, Infrastructure and Physiography	24
5.1	Accessibility	24
5.2	Climate	24
5.3	Local Resources and Infrastructure	24
5.3.1	Background	24
5.3.2	Physical Infrastructure	25
5.4	Adjacent Operations	25
5.5	Physiography, Flora, and Fauna	25
5.5.1	Physiography	25
5.5.2	Flora	25
5.5.3	Fauna	25
6.0	History	27
6.1	Summary	27
6.2	Initial Discovery	27
6.3	Pre-1970 Development	27
6.4	Cordex Syndicate Exploration and Development	27
6.5	Pinson Mining Company Exploration and Development	28
6.6	Homestake-Barrick Exploration	28
6.7	Pinson Project Production Summary	28
6.8	Atna Resources Exploration	30
7.0	Geological Setting	30
7.1	Regional Setting	30
7.2	Pinson Mine Setting	36
8.0	Deposit Type	39
8.1	Sediment-hosted, Carlin-type Gold System	39
9.0	Mineralization	41
9.1	Host Rocks and Structural Environment	41
9.2	Alteration	47
9.3	Mineralized Zone Configuration	48
9.4	Atna’s Phase 2 Program Interpretations	52
9.4.1	Range Front Mineralization	52
9.4.2	Ogee Zone Mineralization	55
10.0	Exploration	57
10.1.	Introduction	57
10.2	Geologic Mapping and Geochemical Sampling	57
10.3	Drilling	57
10.4	Trenching and Channel Sampling	58
10.5	Geophysics	58
10.6	Underground Drifting / Evaluation	59
10.6.1	Atna Underground Drifting / Evaluation	59
11.0	Drilling	61
11.1	Summary of Past and Present Programs	61
11.1.1	Drilling by Earlier Operators	61
11.1.2	Drilling by Atna Resources	61
11.2	Drilling methods	64
11.2.1	Reverse Circulation Rotary Drilling	68
11.2.2	Diamond Core Drilling	68
11.3	Logging	69
11.3.1	Reverse Circulation Rotary Chip Logging	69
11.3.2	Diamond Drill Core Logging	69
12.0	Sampling Method and Approach	70
12.1	Sampling Methods	70
12.1.1	Reverse Circulation Rotary	70
12.1.2	Diamond Core Drilling	70
12.2	Sample Quality – Recovery	70
12.2.1	Reverse Circulation Rotary	70
12.2.2	Core	70
12.3	Sample Interval	71
12.3.1	Reverse Circulation Rotary	71
12.3.2	Core	71
12.4	Sample Preparation, Quality Control Measures and Security	71
12.4.1	Sample Preparation and Quality Control Measures – RC Rotary Drilling	71
12.4.2	Sample Preparation and Quality Control Measures – Core Drilling	72
12.4.3	Security – Reverse Circulation and Core Samples	74
12.4.4	Sample Preparation	74
12.5	Certified Standard Insertion	75
12.5.1	Protocol	75
12.5.2	Summary of Results	76
12.6	Blank Sample Insertion	79
12.6.1	Protocol	79
12.6.2	Summary of Results	79
12.7	Duplicate Samples – Reverse Circulation Rotary	80
12.7.1	Protocol	80
12.7.2	Summary of Results	81
12.8	Check Assays of Mineralized Samples	82
12.8.1	Protocol	82
12.8.2	Summary of Results	83
12.9	Laboratory Quality Assurance and Control	87
12.10	Phase 2 QA/QC	88
12.10.1	Certified Standard Insertion	88
12.10.2	Phase 2 - Blank Sample Insertion	90
12.10.3	Duplicate Samples – Reverse Circulation Rotary	91
12.10.4	Check Assays of Mineralized Samples	93
12.10.5	Laboratory Quality Assurance and Control	94
13.0	Data Verification	96
13.1	Summary	96
13.2	Database of Previous Drilling	96
13.2.1	Table Names	96
13.2.2	Data Corrections	97
13.2.3	General description of pre-Atna assay results and procedures	99
13.3	Pulp Selection	100
13.4	Re-Assay Methods	100
13.4.1	Processing by Inspectorate	100
13.4.2	Processing by ALS Chemex	101
13.5	Re-Assay Results	102
13.6	Discussion	104
14.0	Database Audit	105
14.1	Analytical Data	105
14.1.1	Validation Process – Phase 1 Program	105
14.2	Analytical Data – Phase 2 Program	106
14.2.1	Validation Process – Phase 2 Program	106
14.3	Drill-hole Collar Location	107
14.3.1	Drill-hole Collar Location - Phase 1 Program	107
14.3.2	Drill-hole Collar Location - Phase 2 Program	107
15.0	Adjacent Properties	108
15.1	Summary	108
15.2	Preble Mine	108
15.3	Getchell/Turquoise Ridge Mine Complex	109
15.3.1	History	109
15.3.2	Geologic Setting	109
15.3.3	Gold Geology	110
15.4	Twin Creeks	111
15.4.1	Physiography	111
15.4.2	Stratigraphy	111
15.4.3	Structure	112
15.4.4	Erosion & Alluviation	112
15.4.5	Mineralization	113
15.4.6	Alteration	113
15.4.7	Twin Creeks Summary	114
16.0	Mineral Processing and Metallurgical Testing	115
16.1	Summary of Metallurgical Test Work	115
16.2	Processing Options	119
16.3	Possible Treatment Plants and Potential Costs	119
16.3.1	Treatment Plants	119
16.3.2	Potential Treatment Costs	120
16.4	Impurity Levels and Mineralized Material Types	120
16.4.1	Laboratory Classification of Metallurgical Mineralization Types	121
16.5	Planning and test work – general	123
17.0	Mineral Resources Estimates	124
17.1	March 2005 Resource Calculation	126
17.1.1	Geologic Model, Domains and Coding	126
17.1.1.1	Geologic Model	126
17.1.1.2	Domains and Coding	126
17.2	Available Data	133
17.3	Compositing	136
17.4	Exploratory Data Analysis	136
17.4.1	Basic Statistics by Domain	136
17.4.2	Contact Profiles	137
17.4.3	Histograms and Log-Probability Plots	139
17.4.4	Conclusions and Modeling Implications	145
17.5	Bulk Density Data	145
17.6	Evaluation of Outlier Grades	145
17.7	Variography	146
17.8	Model Setup and Limits	147
17.9	Interpolation Parameters	150
17.10	Validation	151
17.10.1	Visual Inspection	151
17.10.2	Model Checks for Change of Support	151
17.10.3	Comparison of Interpolation Methods	155
17.10.4	Swath Plots	155
17.11	Resource Classification	159
17.12	Economic and Technical Parameters Used for March 2005 Resource Analysis	164
17.13	March 2005 Mineral Resources	166
17.14	2007 Resource Revision	168
17.14.1	Resource Estimation Method	169
17.14.2	Data Analysis	170
17.14.3	Block model parameters	177
17.15	2007 Resource Estimates	178
17.15.1	2007 Range Front Resource Estimate	179
17.15.2	2007 Ogee Zone Resource Estimate	182
17.15.3	CX-West Resource Estimate	185
17.15.4	Comparison of March 2005 Estimate with 2007 Estimate	186
18.0	Other Relevant Data	189
18.1	Potential Mining Methods	189
18.2	Preliminary Geotechnical Evaluations	189
18.3	Potential Drift and Fill Mining Method Layout	190
18.4	Description of Potential Mine Development	193
18.5	Analysis of Ground Support Requirements	194
18.6	Required Backfill & Equipment	194
18.7	Dewatering	195
18.9	Infrastructure	195
18.9.1	Water Supply	195
18.9.2	Power Supply	195
18.9.3	Buildings	196
18.10	Environmental and Socio-Economic	196
18.10.1	Environmental	196
18.10.2	Environmental Permitting	198
18.10.3	Population, Demographics & Ethnicity	199
18.10.4	Employment	199
18.10.5	Workforce Qualifications	199
19.0	Conclusions	200
20.0	Recommendations	201
	References	202
	Appendices	205

FIGURES

Figure 4-1: Location Map of the Pinson Mine Project	14
Figure 4-2: Generalized Land Status Map	15
Figure 4-3: Detailed Land Status and Agreement Map	17
Figure 7-1: Regional Stratigraphy	32
Figure 7-2: Regional Geology	34
Figure 7-3: TectonoStratigraphy	35
Figure 7-4: Mine Site Geology	38
Figure 9-1: Geologic Cross Section 6800NE	44
Figure 9-2: Geologic Cross Section 7100NE	45
Figure 9-3: Geologic Cross Section 7700NE	46
Figure 9-4: Grade-Thickness Long Section, CX Fault Zone	49
Figure 9-5: Grade-Thickness Long Section, Range Front Fault Zone – Phase 1 Results	51
Figure 9-6: Grade-Thickness Long Section, Range Front Fault Zone – Phase 1 & 2	53
Figure 9-7: Grade-Thickness Long Section, Upper Range Front Fault Zone – Phase 1 & 2	54
Figure 9-8: Ogee Zone Grade Shell	56
Figure 12-1: Standards Run With Gravimetric Finish	78
Figure 12-2: Standards Run With AA Finish	78
Figure 12-3: Comparison of Duplicate Samples Taken at the Drill vs. Assay Sample	82
Figure 12-4: Sample Variance of All Check Samples	83
Figure 12-5: X-Y Pair Plot of Check vs. Original Assay, By Grade	84
Figure 12-6: Variance of AA-Finish samples < 3000 ppb	85
Figure 12-7: Gravimetric Finish Comparisons – -Check Assays to Original Assays	86
Figure 12-8: Laboratory Blind Duplicates	87
Figure 12-9: Analytical Standard Sample Results – Phase 2 Program	89
Figure 12-10: RockLabs Moving Average Method – Phase 2 Program	90
Figure 12-11: RC Duplicate Sample comparison – Phase 2 Program	92
Figure 12-12: Check Assay Comparison Graph – Phase 2 Program	94
Figure 12-13: Internal Laboratory Quality Control Samples – Phase 2 Program	95
Figure 13-1: Test Result Variances	103
Figure 17-1: Rotated Vertical NW-SE Cross-Sections	125
Figure 17-2: Cross Sectional View of CX and Range Front Fault Zones	127
Figure 17-3: Isometric View of CX and Range Front Fault Zones	128
Figure 17-4: Sectional View of HG Zone Trends Within Fault Zones	129
Figure 17-5: Comparison of Indicator Probability Estimation Techniques - Plan	130
Figure 17-6: Comparison of Indicator Probability Estimation Techniques - Section	131
Figure 17-7: CX High-Grade Zone Modifications	132
Figure 17-8: Range Front HG Zone Modifications	132
Figure 17-9: Distribution of Drill Holes by Type - CX Zone	134
Figure 17-10: Distribution of Drill Holes by Type – Range Front Zone	135
Figure 17-11: Contact Profile of CX HG vs. LG Domains	138
Figure 17-12: Contact Profile of Range Front HG vs. LG Domains	138
Figure 17-13: Histogram of Gold in CX Fault Zone	139
Figure 17-14: Histogram of Gold in CX HG Zone	140
Figure 17-15: Histogram of Gold in CX LG Portion of Fault Zone	140
Figure 17-16: Histogram of Gold in Range Front Fault Zone	141
Figure 17-17: Histogram of Gold in Range Front HG Zone	141
Figure 17-18: Histogram of Gold in Range Front LG Portion of Fault Zone	142
Figure 17-19: Log-Probability Plot of Gold in CX Fault Zone	142
Figure 17-20: Log-Probability Plot of Gold in CX HG Zone	143
Figure 17-21: Log-Probability Plot of Gold in CX LG Portion of Fault Zone	143
Figure 17-22: Log-Probability Plot of Gold in Range Front Fault Zone	144
Figure 17-23: Log-Probability Plot of Gold in Range Front HG Zone	144
Figure 17-24: Log-Probability Plot of Gold in Range Front LG Portion of Fault Zone	145
Figure 17-25: CX Zone Block Model Limits	148
Figure 15-26: Range Front Zone Block Model Limits	149
Figure 17-27: CX HG Zone Drill Hole and Block Grade Distribution	152
Figure 17-28: Range Front HG Zone Drill Hole and Block Grade Distribution	153
Figure 17-29: CX HG Zone Krige/IDW/Herco Plots	154
Figure 17-30: Range Front HG Zone Krige/IDW/Herco Plots	154
Figure 17-31: Swath Plot CX HG Zone, East-West	156
Figure 17-32: Swath Plots CX HG Zone, North-South	156
Figure 17-33: Swath Plot CX HG Zone, Vertical	157
Figure 17-34: Swath Plot Range Front HG Zone, East-West	157
Figure 17-35: Swath Plot Range Front HG Zone, North-South	158
Figure 17-36: Swath Plot Range Front HG Zone, Vertical	159
Figure 17-37: Confidence Limits Distribution by Drill-hole Spacing	161
Figure 17-38: CX Zone Resource Classification	163
Figure 17-39: Range Front Zone Resource Classification	164
Figure 17-40: Range Front Cummulative Probability	172
Figure 17-41: Range Front Cummulative Frequency Curve	173
Figure 17-42: Ogee Cummulative Probability	174
Figure 17-43: Ogee Cummulative Frequency Curve	175
Figure 17-45: CX-West Cummulative Frequency Curve	177
Figure 17-46: Gold grade distribution in the Upper Range Front zone.	181
Figure 17-47: Gold grade distribution in the Ogee zone	183
Figure 17-48: Gold grade distribution in the Ogee zone	184
Figure 17-49: Grade distribution in CX-West Fault	186
Figure 18-1: Schematic 4500-Level – Range Front and Ogee Zones	191
Figure 18-2: Schematic 4700 Level Production Plan – Range Front and Ogee Zones	192
Figure 18-3: Schematic Production Access Development	193
Figure 18-4: Mine Facilities	197

TABLES

Table 1-1: Pinson Project Mineral Resource Summary	7
Table 1-2: Pinson Project Proposed Program Budget	9
Table 4-1: Royalty Summary	22
Table 5-1: Seasonal Temperature Variations	24
Table 6-1: Pinson Property Production Summary	29
Table 10-1: Ogee Zone Channel Sample Assays	60
Table 11-1: Summary of Drilling	61
Table 11-2: Summary of Atna Resources Phase I Drilling	64
Table 11-3: Summary of Atna Resources Phase 2 Surface Drilling	65
Table 11-4: Summary of Atna Resources Phase 2 Underground Diamond Drilling	66
Table 12-1: Phase 1 - RockLabs Reference Material	75
Table 12-2: Phase 1-Failed Gravimetric standards	77
Table 12-3: Decorative Stone (Blank Sample) Analysis	80
Table 12-4: Statistics of Lab Variance in AA vs. Gravimetric Finish	84
Table 12-5: Failure Rate for 3 Check Assay Jobs	86
Table 12-6: Standards used in Phase 2	88
Table 12-7: List of Standard Assay Failures. Phase 2 drill program	88
Table 12-8: Blank Sample Assay Failures. Phase 2 drill program	91
Table 12-9: Duplicate RC Samples Beyond Acceptable Limits-Phase 2 program	92
Table 12-10: Check Assay Results Beyond Acceptable Limits-Phase 2 program	93
Table 13-1: Field Definitions in (HMC) Pinson Database	98
Table 13-2: CDN Resource Laboratory Standards	101
Table 13-3: Analyses of Standards Used in Testing	103
Table 16-1: Composites by Zone	116
Table 16-2: Summary of Metallurgical Assay Results	117
Table 16-3: Results of Autoclave - Leach Procedure	118
Table 16-4: Metallurgical Impurity Levels (Preliminary Information)	121
Table 16-5: Preliminary Summary of Classification of Mineralized Material	122
Table 17-1: Summary of Geologic Domains	133
Table 17-2: Drill-hole Sample Statistics by Fault Zone	137
Table 17-3: Drill-hole Sample Statistics by HG/LG Portion of Fault Zone	137
Table 17-4: Proportion of Contained Gold in Decile/Percentile of Samples	146
Table 17-5: Variogram Parameters	147
Table 17-6: CX Zone Block Model Limits	149
Table 17-7: Range Front Zone Block Model Limits	149
Table 17-8: Interpolation Parameters for Ordinary Krige Estimates	150
Table 17-9: Interpolation Parameters for IDW Estimates	151
Table 17-10: Comparison of Interpolation Methods	155
Table 17-11: Quarterly and Yearly Confidence Limits Determination	160
Table 17-12: March 2005 Mineral Resource, CX Zone	166
Table 17-13: 2005 Mineral Resource, Range Front Zone	167
Table 17-14: 2005 Mineral Resource, Combined CX and Range Front Zones	168
Table 17-15: Drill-hole Sample Statistics by Fault Zone	171
Table 17-16: Block Model Parameters	178
Table 17-17: Search Ellipse Parameters	178
Table 17-18: 2007 Revised Range Front Resource Estimate	180
Table 17-19: Ogee Zone Resource Summary	182
Table 17-20: CX-West Resource Estimate	185
Table 17-21: Combined Resource Summary (RF, CX, Ogee, CX-West)	187
Table 17-22: March/2005 versus 2007 Resource-Upper Range Front	188
Table 18-1: Key Permits	198
Table 18-2: Other Required Permits for Mining	199

Figure 17-20: Log-Probability Plot of Gold in CX HG Zone

Figure 17-21: Log-Probability Plot of Gold in CX LG Portion of Fault Zone

135

Figure 17-22: Log-Probability Plot of Gold in Range Front Fault Zone

Figure 17-23: Log-Probability Plot of Gold in Range Front HG Zone

136

Figure 17-24: Log-Probability Plot of Gold in Range Front LG Portion of Fault Zone

17.4.4 Conclusions and Modeling Implications

The results of the EDA indicate that the HG zones in the CX and Range Front fault zones are truly distinct in nature and should be isolated from the lower-grade material during the grade estimation process. This conclusion is best supported by the results of the contact profile analysis and the log-probability distributions of the data.

17.5 Bulk Density Data

There are no individual bulk density values available in the drill-hole database from which to conduct interpolation estimations in the block model. Historical data derived from the production experience at the Pinson mine indicates an average bulk density tonnage factor of 13 cubic feet per ton of material. This value has been retained for this resource estimation.

17.6 Evaluation of Outlier Grades

The presence of extreme sample grades was evaluated on the histograms and log-probability plots shown in Figures 17-11 through 17-21. There are few indications of anomalous values other than a few data points in the upper grade range of the Range Front LG portion of the fault zone (a domain which is not included in the final resource summary).

A decile analysis of the data was also conducted in order to identify the possible existence of anomalous values. If the top-decile of the database contains more than 40% of the contained gold, or there is more than twice the contained gold than the previous (9th) decile, then some form of top-cutting may be required and the data must then be evaluated on a finer (percentile) scale. At this stage, if there is>10% of the contained gold in a single percentile bin, or there is more than twice the contained gold than the previous bin, then some form of top-cutting may be

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required. The proportion of gold in the HG and LG domains of the CX and Range Front zones is summarized in table 17-4.

Table 17-4: Proportion of Contained Gold in Decile/Percentile of Samples

	Percent of contained gold (%)
*Decile/Percentile*	*CX HG zone*	*CX LG Zone*	*RF HG Zone*	*RF LG Zone*
80	15.2	18.8	15.1	16.0
90	31.4	68.3	32.0	73.1
98	4.4	9.1	4.7	10.2
99	5.3	21.3	5.3	23.0

The results in the table above show that there are no anomalous values in the two HG domains. There are 7 samples in the CX HG zone which are greater than 1 opt Au, 5 of which are less than 1.1 opt Au. The other two samples (at 1.82 and 1.84 opt Au) are from drill hole RHA-1567 and are in a thick area of mineralization with lots of proximal samples. As a result, the overall effect of these two high-grade samples on the resource is minimal.

The Range Front HG zone has 5 samples which exceed 1 opt Au. These samples were found to be supported by adequate additional proximal samples and it is felt that no action with respect to top-cutting is required.

Both LG domains contain several high-grade samples which are considered anomalous. As stated previously, these are relatively high-grade samples which do not conform to the interpretation of the HG domain. Based on the current understanding of the nature of the HG zone, these intervals are probably interconnected with the main HG portion of the deposit but this interpretation cannot be supported by the current density of drilling. Therefore, these few intervals are essentially excluded from the resource estimate and are assigned to the LG zone. The LG composite samples have been top-cut to a value of 0.25 opt Au. This involves a total of 9 samples in the CX LG zone and 5 samples in the Range Front LG zone.

17.7 Variography

The degree of spatial variability in a mineral deposit depends on both the distance and direction between points of comparison. Typically, the variability between samples increases as the distance between samples also increases. If the degree of variability is related to the direction of comparison, then the deposit is said to exhibit anisotropic tendencies which can be summarized with the search ellipse. The semi-variogram is a common function used to measure the spatial variability within a deposit.

The components of the variogram include the nugget, sill and range. Often, samples compared over very short distances (even samples compared from the same location) show some degree of variability. As a result, the curve of the variogram often begins at some point on the y-axis above the origin – this point is called the “nugget”. The nugget is a measure of not only the natural variability of the data over very short distances but also a measure of the variability that can be introduced due to errors during sample collection, preparation and assaying.

The degree of variability between samples typically increases as the distance between the samples becomes greater. Eventually, the degree of variability between samples reaches a

138

maximum value. This is called the “sill” and the distance between samples at which this occurs is referred to as the “range”.

The spatial evaluation of the data in this report has been conducted using a correlogram rather than the traditional variogram. The correlogram is normalized to the variance of the data and is less sensitive to outlier values, generally giving better results.

Correlograms have been produced from the 5-foot composite drill-hole sample data for the CX HG and Range Front HG domains using the commercial software program Sage 2001 (Isaacs and Co). Multidirectional correlograms were generated at 30o intervals both horizontally and vertically resulting in a total of 37 sample correlograms in which an algorithm determines the best-fit model. The sample correlograms are summarized in Table 17-5.

17.8 Model Setup and Limits

Although the CX and Range Front zones are only separated by about 600 feet, two separate block models were initialized in order to minimize their size and thus, reduce calculation and manipulation times. Both models utilize the same block size (10x10x10 ft) and both models have been rotated by 30 degrees which approximately aligns it with the general trend of the zones and reduces the total number of blocks in the model.

The limits and orientation of the CX block model are shown in Figure 17-25.

Table 17-5: Variogram Parameters

Domain	Nugget	S1	S2	1^st Structure			2^nd Structure
Domain	Nugget	S1	S2	Range (ft)	AZ	Dip	Range (ft)	AZ	Dip
CX HG	0.550	0.181	0.269	461	4	-3	837	180	22
				157	96	-32	571	73	37
				9	89	58	40	113	-45
RF HG	0.500	0.483	0.015	195	94	22	688	8	23
				146	203	38	626	227	61
				8	341	44	305	106	16

139

Figure 17-25: CX Zone Block Model Limits

The CX zone model limits are summarized in Table 17-6.

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Table 17-6: CX Zone Block Model Limits

*Direction*	*Min*	*Max*	*Size (ft)*	*#Blocks*
X	0	1800	10	180
Y	0	2600	10	260
Z	3400	5200	10	180

(Origin 8500E, 9500N rotated 30 degrees from north)

The limits and orientation of the Range Front block model are shown in Figure 17-26 and are summarized in Table 17-7

Figure 17-26: Range Front Zone Block Model Limits

Table 17-7: Range Front Zone Block Model Limits

*Direction*	*Min*	*Max*	*Size (ft)*	*#Blocks*
X	0	2200	10	220
Y	0	2700	10	270
Z	2600	5600	10	300

(Origin 8100E, 11200N rotated 30 degrees from north)

17.9 Interpolation Parameters

The block model grade interpolation, by ordinary kriging (OK), was conducted using hard-boundary code matching within the Zone domains. This means that the block grade estimations within each domain were restricted to drill-hole samples located within that domain.

The results of the OK estimation were compared with the Hermitian (Herco) polynomial change of support model (also referred to as the Discrete Gaussian correction, this method is described in detail in Section 17.11.2) utilizing a series of pseudo grade/tonnage distribution curves. Modifications were made to the krige interpolation parameters until the desired results were obtained.

The CX and Range Front OK models have been generated with a relatively limited number samples in order to match the Herco distribution. This approach reduces the amount of smoothing (averaging) in the model and, while there may be some uncertainty on a localized scale, this approach produces reliable estimations of the recoverable grade and tonnage for the overall deposit.

The interpolation in each zone was conducted using a search ellipse oriented sub-parallel to the general trend of the mineralization. The CX zone uses a search ellipse measuring 700x700x100 feet oriented at an azimuth of 40 degrees and dips –60 degrees SE. The Range Front zone uses an ellipse measuring 900x900x100 feet at an azimuth of 15 degrees and –65 SE dip.

The interpolation parameters used for the OK estimate are summarized in Table 17-8.

Table 17-8: Interpolation Parameters for Ordinary Krige Estimates

Domain	Search						# Comps
	Range (ft)			Orientation (AZ,Dip)			Max/Min per Blk/
	X	Y	Z	X	Y	Z	Max per hole
CX HG	700	700	100	130,-60	40,0	310,-30	6/3/2
RF HG	900	900	100	105,-65	15,0	285,-25	12/4/3

For comparison purposes, grade estimates were also conducted using both the inverse distance (IDW) interpolation method and a nearest neighbor (NN) distribution. Note that the IDW estimates were also “tuned” in comparison with the Herco grade/tonnage distribution.

As stated previously, the resource estimation is summarized within the HG zone domains. However, the gold content in the low-grade portion of the fault zones has also been estimated for the purposes of dilution studies in future mine design evaluations. The LG zone estimations were made using the relative elevation IDW method used in the indicator probability estimates described in section 17.2.2 of this report.

The IDW interpolation parameters are summarized in the table below.

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Table 17-9: Interpolation Parameters for IDW Estimates

Domain	Search						# Comps
	Range (ft)			Orientation (AZ,Dip)			Max/Min per Blk/
	X	Y	Z	X	Y	Z	Max per hole
CX HG	700	700	100	130,-60	40,0	310,-30	16/5/4
RF HG	900	900	100	105,-65	15,0	285,-25	30/6/5
CX LG	700	700	700	Relative to gold zone trend			20/1/4
RF LG	900	900	900	Relative to gold zone trend			20/1/4

17.10 Validation

The resource model results were validated in several ways including a thorough visual review of the results, comparisons with the change of support model, comparisons with other methods and grade distribution comparisons using swath plots.

17.10.1 Visual Inspection

Detailed visual inspection of the block models has been conducted in both cross section and plan. This includes the proper coding and percentage of blocks with respect to the respective domains. The distribution of block grades were also compared relative to the drill-hole samples in order to ensure proper representation in the model. Examples of cross sections showing the drill-hole and block model grade distributions are shown in Figure 17-25 for the CX zone and Figure 17-26 for the Range Front zone. Complete sets of cross sections through the models are included in Appendix 17-3 of the December 2005, Sedar filed, 43-101 Technical Report.

17.10.2 Model Checks for Change of Support

The relative degree of smoothing in the block model estimates were evaluated using the Discrete Gaussian of Hermitian Polynomial Change of Support method (described by Journel and Huijbregts, Mining Geostatistics, 1978). With this method, the distribution of the hypothetical block grades can be directly compared to the estimated (OK and IDW) model through the use of pseudo-grade/tonnage curves. Adjustments are made to the block model interpolation parameters until an acceptable match is made with the Herco distribution. In general, the estimated model should be between 5-10% higher in tonnage and 5-10% lower in grade when compared to the Herco distribution at the projected cut-off grade. These differences account for selectivity and other potential ore-handling issues which commonly occur during mining.

The Herco distribution is derived from the declustered composite grades that have been adjusted to account for the change in support as one goes from smaller drill-hole composite samples to the large blocks in the model. The transformation results in a less skewed distribution but with the same mean as the original declustered samples. Note that the standardized block variance values resulting from the correlogram-based OK estimate are multiplied by the variance of the declustered composite data to obtain relative block variances used in the Herco analysis.

Comparisons between the krige/IDW and Herco models are shown for the CX and Range Front HG zones in Figures 17-27 and 17-28.

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Figure 17-27: CX HG Zone Drill Hole and Block Grade Distribution

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Figure 17-28: Range Front HG Zone Drill Hole and Block Grade Distribution

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Figure 17-29: CX HG Zone Krige/IDW/Herco Plots

Figure 17-30: Range Front HG Zone Krige/IDW/Herco Plots

The results show good correlation at a typical cut-off limit for an underground operation of between 0.2 and 0.3 opt Au. The separation of the lines at higher cut-offs are an indication that the OK model may underestimate the higher-grade portions of the deposit and, conversely, the IDW model may slightly overestimate the higher grades.

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17.10.3 Comparison of Interpolation Methods

The OK, IDW and NN models are tabulated for comparison purposes in Table 15-10.

Table 17-10: Comparison of Interpolation Methods

*Cut-off*	*OK Model*		*IDW Model*		*NN Model*
(Auopt)	kTons	Au opt	kTons	Au opt	kTons	Au opt
CX Zone :
0	2,052	0.25	2,052	0.24	2,052	0.24
0.15	1,718	0.27	1,744	0.26	1,149	0.36
0.2	1,404	0.29	1,242	0.29	880	0.41
0.25	982	0.32	658	0.35	687	0.46
0.3	419	0.39	397	0.40	514	0.53
RF Zone :
0	4,298	0.32	4,298	0.33	4,298	0.31
0.15	4,251	0.33	4,215	0.34	3,167	0.38
0.2	3,952	0.34	3,903	0.35	2,499	0.44
0.25	3,145	0.36	3,157	0.38	2,099	0.48
0.3	1,908	0.42	2,131	0.43	1,539	0.55

The variance on gold grades in the OK and IDW models is less generally than 5% at all cut-off limits. Local variances in tonnage and grade occur primarily at depth, where the drill-hole spacing increases.

17.10.4 Swath Plots

A swath plot is a graphical display of the grade distribution derived from a series of bands, or swaths, generated in several directions through the deposit. The grade variations from the OK and IDW models are compared using the swath plot to the distribution derived from the declustered (NN) grade model.

On a local scale, the NN model does not provide reliable estimations of grade but, on a much larger scale, it represents an unbiased estimation of the grade distribution based on the underlying data. Therefore, if the OK and IDW models are considered unbiased, their grade trends may show local fluctuations on a swath plot but, the overall trend should be similar to the NN distribution of grade.

Swath plots have been generated for the gold grade distribution in both the CX HG zone, shown in Figures 17-31, 17-32 and 17-33 and the Range Front HG zone shown in Figures 17-34, 17-35 and 17-36. It should be noted that the 50-foot wide swaths have been applied to the rotated block model. Therefore, the X-axis increments in the plots are listed by swath number, rather than actual mine grid coordinate values.

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Figure 17-31: Swath Plot CX HG Zone, East-West

Figure 17-32: Swath Plots CX HG Zone, North-South

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Figure 17-33: Swath Plot CX HG Zone, Vertical

Figure 17-34: Swath Plot Range Front HG Zone, East-West

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Figure 17-35: Swath Plot Range Front HG Zone, North-South

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Figure 17-36: Swath Plot Range Front HG Zone, Vertical

There are local fluctuations which tend to increase in intensity at depth or along the fringes of the deposits where the drilling spacing tends to increase. In general, the results are similar between the models with no indications of significant grade bias.

17.11 Resource Classification

A common method used in the classification of mineral resources involves geostatistical methods that define categories based on the confidence limits of the estimation. Measured resources are defined as material in which the predicted grades are within ±15% accuracy on a quarterly basis, at a 90% confidence limit. In other words, there is a 90% chance that the predicted grades for a quarter-year of production will be within ±15% of the actually achieved production grades. Similarly, Indicated resources include material in which the yearly production grades are estimated with ±15% accuracy at the 90% confidence limit.

The method is based on the large sample normal theory that assumes that as the grade estimations from smaller blocks are combined into larger ones, the errors of the estimation become normally distributed (as described by B. Davis, 1997). The steps in generating the classification parameters for the CX and Range Front zones are described as follows.

This exercise assumes a nominal daily production rate from either of the zones of 500 tons per day which equates to a monthly production rate of approximately 15,000 tons per month. At an average tonnage factor of 13 cubic feet per ton, a block measuring 60x60x60 feet represents approximately one month of production (16,600 tons).

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A block equal in size to the volume of one month’s production is created and the kriging variance is determined using a series of theoretical drill holes at intervals averaging 50, 100 and 200 foot spacing. The calculations are done over a series of drill-hole grids in order to evaluate the variation in the results with respect to the spacing of the drill data.

The correlogram used to determine the kriging variance in the large block is derived from the actual gold sample data which has been composited to 5-foot intervals (Table 15-5). Because the correlogram was used, the normalized block kriging variance (a variable which is output from the OK run) was standardized to the underlying data by multiplying by the square of the coefficient of variation (CV=std dev/mean from the original 5 ft composite data).

The relative standard error for a quarter year of production is determined by taking the square root of the standard block variance divided by three (i.e. divide by 3 for a quarter-yr of production or divide by 12 in order to determine the error for a full year of production). Finally, the 90% confidence limit is determined by multiplying the relative standard error by 1.645. The results of the exercise are listed in Table 16-11 and shown in graphical form in Figure 17-37.

Table 17-11: Quarterly and Yearly Confidence Limits Determination

DH	Norm OK		Std.	Relative Std. Error		Conf. @ 90% Limit
Spacing (ft)	Blk.Var.	(CV)sqd	Blk.Var.	Qtr	Year	Qtr	Year
CX Zone:
200	0.1569	0.833	0.1307	0.2088	0.1044	34.3%	15.2%
100	0.0546	0.833	0.0455	0.1233	0.0616	20.3%	10.1%
50	0.0409	0.833	0.0341	0.1068	0.0533	15.6%	8.8%
Range Front Zone:
200	0.1035	0.886	0.0915	0.1549	0.0874	28.8%	14.4%
100	0.0761	0.886	0.0674	0.1499	0.0750	24.7%	12.3%
50	0.0371	0.886	0.0329	0.1047	0.0523	15.2%	8.6%

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Figure 17-37: Confidence Limits Distribution by Drill-hole Spacing

The results of the exercise are very similar for both the CX and Range Front zones. A quarter of a year of production can be estimated at +/-15% accuracy at the 90% confidence limit with drill holes spaced at approximately 35-foot intervals (projected from the trends exhibited in Figure 17-37). Similarly, one year of production can be estimated at the same confidence levels with drill holes spaced at 175-200 foot intervals. (Note that an even grid of drill holes spaced at 200- foot intervals would result in the maximum distance from a block in the model to a drill-hole sample location of 125 feet).

Some additional factors that were considered in the designation of classification parameters are summarized as follows:

The Pinson deposits are geologically similar to others in the area and the mode of formation is well understood. The nearby Getchell deposit base its classification on the spacing of drill holes. Measured resources vary (by zone) from between 50-100 feet. Indicated resources are defined by drill holes spaced at 150-foot intervals.

·	The CX Zone was in operation and is currently exposed on surface in the walls and floor of the CX pit. This exposure provides information regarding the nature and continuity of the zone and greatly increases the degree of confidence in this portion of the deposit.

Ultimately, the classification definitions are based on the confidence in the continuity of the mineralization and are expressed based on the distance and proximity to the drill-hole sample

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locations. The upper portion of the CX zone has been upgraded to a higher classification due to its exposure on surface.

Typically, a strict numerical classification based on “distance to” protocols can produce minor patches or islands of blocks within large continuous zones of a different class. Retaining these “artifacts” inflicts a certain degree of clutter and confusion when visually reviewing the distribution of resource classes in the model. Secondly, the locally complex class designation gives an impression of detail or complexity in the designation which is often not supported by the current density of drilling. Therefore, some local generalizations were made – manually re-classified in order to produce a cleaner and more consistent class distribution. These manual changes are described as follows. The CX zone “measured” criteria (3 holes within an average distance of 75 feet) included zones which extended below 300 feet below surface – these areas were reclassified (downgraded) as indicated material. The original contact between indicated and inferred material in the Range Front zone was somewhat ragged in places and was more clearly defined through some minor manual reclassification. Ultimately, these changes are considered relatively minor and are implemented to make the various class designations more clearly defined and more visually apparent.

The resource classification parameters are defined as follows.

Measured Resources: Material located within the high-grade domain which meets the two criteria as follows: Blocks which have been estimated by a minimum of three drill holes within an average distance of 75 feet and, second, which are within a maximum distance of 300 feet from surface exposures created during previous production activities. This includes only material in the vicinity of the CX pit.

Indicated Resources: Blocks which do not meet the criteria for measured resources but include material located within the high-grade zone in which grades have been estimated by a minimum of three drill holes within a maximum average distance of 125 feet.

Inferred Resources: The remaining blocks in the high-grade zone which do not meet the criteria for measured or indicated resources. Internally, this is based on drill holes with a maximum spacing of 400 feet in the CX zone and up to 600 feet in the Range Front zone. On the peripheries of the deposit, which is not “closed” by drilling, the resource extends a maximum distance of 150 feet from a drill hole.

The distribution of resources by class in the CX and Range Front zones is shown in Figures 17-38 and 17-39.

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Figure 17-38: CX Zone Resource Classification

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Figure 17-39: Range Front Zone Resource Classification

17.12 Economic and Technical Parameters Used for March 2005 Resource Analysis

For purposes of the March 2005 resource analysis, generalized economic and technical parameters have been estimated for the Pinson project in order to project the relative scale of the potential economic viability of a potential operation and it’s contained mineral resources. It is important to recognize that this is a series of initial assumptions and does not represent an economic analysis for this project. As a result, this report contains only mineral resources and does not include mineral reserves for the Pinson deposits.

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The key economic and technical assumptions, parameters and methods used in this report are primarily derived from proximal existing operations in the Pinson area. Many of the mines in Northern Nevada exhibit similar gold grades and mineralized material type, exploitation methods and metallurgical characteristics. The key assumptions are listed as follows:

1.	Underground mining utilizing underhand, drift-and-fill exploitation methods. Cost estimated at US$50/ton of ore mined (inclusive of backfill costs). These costs are based upon the Company’s initial negotiations of contract rates for mining and development work with Small Mine Development (“SMD”).

2.	Access to the mineralized zones will be via ramp/decline beginning from the bottom of the existing CX-pit floor.

3.	The daily production rate from the combined Range Front and CX zones is estimated to b e 700 tons per day.

4.	Carlin-type, refractory gold ore (similar to Getchell, Meikle, Jerritt Canyon, Deep Star mines), with recoveries estimated to be 93%.

Processed by toll milling at third-party mill (Newmont’s Twin Creeks, Gold Quarry, or Lone Tree mills; Barrick’s Goldstrike complex; and/or Queenstake’s Jerritt Canyon mill). Toll milling costs are estimated to be between $23 and $25/ton milled, including transportation. This is based on initial discussions between Atna and the third-party operators and assumes that 100% of Pinson ore will require either autoclave or roaster processing.

6.	Site indirect and administrative costs (General and Administrative costs) are estimated to be approximately $7/ton.

7.	Projected gold price of $400/oz.

The “base case” cut-off grade is calculated as follows:

Mining ($50/t) + Milling ($25/t) + G&A ($7/t) = total operating cost of $82/ton

Gold price of $400/oz / 34.286 = $11.67/gram

Total Cost / gold price: $82/t / ($400/34.286)

Recovery: (93/100)

----------------------------

=7.56g, or 0.22optAu

(Base case cut-off rounded to 0.20opt Au in tables)

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17.13 March 2005 Mineral Resources

The mineral resources are summarized for the CX zone in Table 17-12, the Range Front zone in Table 17-13 and for both zone combined in Table 17-14.

Table 17-12: March 2005 Mineral Resource, CX Zone

*Category*	*Cut-off* *(Au opt)*	*Tons (000)*	*Grade* *(Au opt)*	*Contained* *Au (koz)*
Measured	0.15	445	0.27	119
	0.20	319	0.30	97
	0.25	213	0.34	73
	0.30	130	0.39	50
Indicated	0.15	502	0.28	143
	0.20	427	0.30	130
	0.25	278	0.34	96
	0.30	156	0.40	63
Measured + Indicated	0.15	947	0.28	262
	0.20	746	0.30	226
	0.25	490	0.34	169
	0.30	286	0.40	113

Inferred	0.15	770	0.27	205
	0.20	658	0.28	185
	0.25	491	0.30	148
	0.30	134	0.36	48

(Base case cut-off grade = 0.20opt Au)

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Table 17-13: 2005 Mineral Resource, Range Front Zone

*Category*	*Cut-off* *(Au opt)*	*Tons (000)*	*Grade* *(Au opt)*	*Contained* *Au (koz)*
Measured	0.15	0	0	0
	0.20	0	0	0
	0.25	0	0	0
	0.30	0	0	0
Indicated	0.15	811	0.32	257
	0.20	721	0.33	241
	0.25	582	0.36	210
	0.30	429	0.39	167
Measured + Indicated	0.15	811	0.32	257
	0.20	721	0.33	241
	0.25	582	0.36	210
	0.30	429	0.39	167

Inferred	0.15	3,440	0.33	1,127
	0.20	3,231	0.34	1,088
	0.25	2,562	0.37	936
	0.30	1,479	0.43	642

(Base case cut-off grade = 0.20opt Au)

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Table 17-14: 2005 Mineral Resource, Combined CX and Range Front Zones

*Category*	*Cut-off* *(Au opt)*	*Tons (000)*	*Grade* *(Au opt)*	*Contained Au (koz)*
Measured	0.15	445	0.27	119
	0.20	319	0.30	97
	0.25	213	0.34	73
	0.30	130	0.39	50
Indicated	0.15	1,313	0.30	400
	0.20	1,148	0.32	371
	0.25	860	0.36	305
	0.30	585	0.39	230
Measured + Indicated	0.15	1,758	0.30	519
	0.20	1,467	0.32	467
	0.25	1,073	0.35	379
	0.30	715	0.39	281

Inferred	0.15	4,211	0.32	1,332
	0.20	3,889	0.33	1,273
	0.25	3,054	0.36	1,084
	0.30	1,612	0.43	690

(Base case cut-off grade = 0.20opt Au)

17.14 2007 Resource Revision

From February to May 2007, resource estimates for the Range Front and Ogee zones were revised and estimates calculated for the CX-West fault zone utilizing both surface and underground drill results obtained after completion of the March 2005 calculation and through the end of Atna’s Phase 2 program in June 2006. This revision includes all analytical results from the Phase 1 and 2 programs as well as all data available from previous operators. Atna completed an additional 56 surface drill holes during Phase 2, including reverse circulation rotary holes (RC), RC pre-collared holes with diamond core-tails, and holes drilled entirely with diamond drilling. Underground diamond drilling included 7 holes drilled in the Range Front zone, 35 holes completed in the Ogee zone, and 6 holes in the CX-West fault zone. Complete details of the Phase 2 drilling are discussed in Section 11 of this report.

At the time of Barrick Gold’s decision to exercise their back in rights (April, 2006), several underground drill holes were in progress targeting the Range Front and CX-West fault zones, and assays from numerous holes from the CX-West, Range Front, and Ogee zones were pending. This additional information was received after the announcement in an April 2006 press release that contained a revision to the March 2005 estimate discussed in section 17.1 to

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17.13 of this report. Geologic interpretations regarding distribution of gold grade and structures in the hanginwall of the upper range front zone have also changed significantly since the March 2005 estimate. This new information requires a revision to the resource estimates made earlier in the Upper Range Front, and Ogee zones, and an estimate to be created on the CX-West fault zone.

Beccause the deepest drill hole on the Range Front zone during Atna’s Phase 2 program intersected the fault above the 4190-foot elevation, no revision has been made in the lower Range Front zone below the 4200-foot elevation.

17.14.1 Resource Estimation Method

The commercial software package Surpac Vision 5.2c, was utilized in all resource revisions and calculations in 2007. Methodologies followed standard practice for generating a geologic based estimate and included the following steps;

The generation of two different oriented, 1:1200 scale, cross section sets with interpreted geology and mineralization (grade shell domains) were constructed from drill intercepts at a 0.1 opt cut. A 0.10 opt cut was chosen as it represents greater than the 90^th percentile (0.072 opt) of all assays contained in the three zone domains, and it was roughly half the value of the proposed 0.2 opt cut based on preliminary economics as outlined in section 16.

2.	Digitizing and 3D modeling of both cross section sets with comparison and modifications made to generate a single, uniform 3 dimensional model of the geology and mineralization.

3.	Slicing of the geologic and mineralization models into 30-foot bench level elevation plans. This step checked the accuracy of section interpretations, with adjustments made to geologic and mineralization models where necessary.

Grade shells (0.1 oz/ton or above) were further modified at the plan level to meet the following criteria: Grade shell boundaries were drawn halfway between drill holes meeting the 0.1 opt cutoff and those that were either barren, or contained intercepts at less than the 0.1 opt cutoff. This rule applied to drill spacings of 600 feet or less. In areas where drill densities were greater than 600 feet, or where margins of the grade shells could not be constrained by drilling, grade shells were drawn at no greater than 150 feet from the closest drill hole.

The purpose of utilizing grade shells is to constrain geologically the interpolation by Surpac to the higher-grade portions of the fault zones where confidence in the geologic continuity of mineralization was high. This approach, rather than a statistical approach, was taken because drill density and underground geology exposures has led to greater understanding on mineralizing controls than was available in previous in the March 2005 estimates. The grade shells apply constraints on the software in two different ways. First, drill intercepts for interpolation are hard coded. That is, drill hole intercepts spearing the grade shell solid were extracted from the data base using a 3DM (solid) - drill hole intersection routine. Drill hole assays were then extracted from these intercepts and composited to 5-foot lengths. The result being a series of 5 foot assay samples contained entirely within the grade shell for each drill hole intersecting the solid. Secondly, the grade shells form a boundary preventing the estimation algorithm from utilizing samples outside of the shell for interpolating grade of blocks within the shell. These grade shells can be considered as “domains” and were used to prevent the

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program from incorporating assay samples which would not be representative of mineralization occurring within a particular domain.

5.	Drill hole intercepts were extracted based on the grade shells and assays for each hole were composited using a down hole method on 5-foot widths.

A preliminary data analysis was completed on the composited assays for each grade shell. This procedure is done to identify multiple populations due to different structural orientations, and outlier grade values. Although it is not truly desirable to cap outlier grades, those outlier grades may, if not sufficiently constrained by drilling, tend to overestimate tons and grade.

7.	A single Block model was constructed covering the three zones (Range Front, Ogee, and CX-West) being estimated. Each zone was, however, inbterpolated separately using the grade shells as constraints and using unique search parameters as guided by structural orientation within the zones.

The method used for the 2007 resource estimation to interpolate block grades for all three zones is inverse distance weighted (IDW) to the 3^rd power. Surpac uses a 3 axis search ellipsoid for all estimation methods that modifies the search based on bearing, dip, plunge, and modifies distance weights for samples differently depending on anisotropy ratios of the search ellipse.

17.14.2 Data Analysis

Statistical evaluation of the composited assay data was completed for each grade shell zone to determine basic data characteristics, evaluate potential for multiple populations of data, and determination of outlier values as outlined in Sections 17-4 and 17-6. Determination of multiple populations and outliers is critical in the Range Front and Ogee zones where structures of differing orientations control mineralization. Summary statistics and data evaluations are provided separately for each grade shell (domain) in Table 17-15, and Figures 17-40 to 17-45. For purposes of data evaluation, all less than detection limit gold values were converted to 0.0001, and all missing intervals where no assay was taken were handled as null values (ie. ignored – as if no samples were taken) in the program.

The larger sample sizes obtained from the Ogee zone and the CX-West zone are due to the inclusion of blast hole data that exists for mineralization, as defined by drilling, past production, and geology, that can be incorporated into the grade shells for both zones. The blast hole data was used to help interpolate grade from prior production into the upper reaches of the CX-West and Ogee mineralized zones at and below pit level. Grade estimates for blocks above the bottom pit levels were eliminated and have not been included in the grade and tonnage calculations.

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Table 17-15: Drill-hole Sample Statistics by Fault Zone

*Gold (opt)*	*Range Front Fault*	*Ogee Zone*	*CX-West Fault*
# samples	421	823	913
Min	0.0001	0.0001	0.0001
Max	2.479	3.036	0.909
Mean	0.309	0.364	0.212
Std Dev	0.310	0.415	0.181
90^th percentile	0.743	0.790	0.486

(Statistics based on 5-foot drill-hole composite samples.)

17.14.2.1 Range Front zone data analysis

Figure 17-40 is a histogram of gold grades with a cumulative probability curve superimposed for the Range Front grade shell. Basic statistics for figure 17-40 are given in Table 17-15. The distribution of grades is highly skewed with potential outliers above 1.5 opt Au. Raw data for the range front zone shows most gold values within the grade shell to fall between 0.1 and 0.48 opt (80^th percentile), with about 18% of the assays occurring above 0.55 opt. Figure 17-41 shows a histogram of the same values that have been transformed using natural log function. In Figure 17-41, the histogram shows a normal histogram curve, while the cumulative frequency curve shows a typical shape for a set of samples coming from a single population. Potential outliers occur above 1.25 opt Au in this histogram. As outlined in Section 17.6, a decile analysis shows 32% of the gold contained in the Range Front grade shell occurs above the 90^th percentile, with 20% of the gold occurring in assays from the 80^th percentile (0.480). Visual inspection of the intercepts show a uniform and widespread distribution in all grade ranges, including those above 1.0 opt, and sufficient assays and drilling surrounding the high grade assays (>1.0 opt Au) support the higher grade zones. The statistical analyses and visual observations indicate that no top cutting be required for the range front zone.

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Figure 17-40: Range Front Cummulative Probability

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Figure 17-41: Range Front Cummulative Frequency Curve

17.14.2.2 Ogee zone data analysis

Figure 17-42 is a histogram of gold grades with a cumulative probability curve superimposed for the Ogee grade shell. Basic satistics for Figure 17-42 are given in Table 17-15. The distribution of raw gold grades is highly skewed with potential outliers above 1.15 opt Au. Raw data for the Ogee zone shows most gold values within the grade shell to fall below 0.500 opt (3^rd quartile). Figure 17-43 shows a histogram of the same values that have been transformed using natural log function. In Figure 17-43, the histogram shows a normal histogram curve, while the cumulative frequency curve shows a typical shape for a set of samples coming from a single population. Potential outliers occur above 1.5 opt Au in this histogram As outlined in Section 17.6, a decile analysis shows 38% of the gold contained in the range front grade shell occurs above the 90^th percentile (0.790 opt Au), with 16% of the gold occurring in assays from the 80^th percentile (0.551 opt Au). Visual inspection of the intercepts show a uniform and widespread distribution in all grade ranges, including those above 1.0 opt, and with sufficient drilling surrounding the high grade assays (>1.0 opt Au) to support and constrain the interpolation of higher grades. The statistical analyses and visual observations indicate that no top cutting be required for the Ogee zone.

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Figure 17-42: Ogee Cummulative Probability

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Figure 17-43: Ogee Cummulative Frequency Curve

17.14.2.3 CX-West zone data analysis

Figure 17-44 is a histogram of gold grades with a cumulative probability curve superimposed for the CX-West grade shell. Basic statistics for Figure 17-44 are given in Table 17-15. The distribution of raw gold grades is moderately skewed with a more normal distribution apparent in the raw data than for both Range Front and Ogee zones. Grade for the CX-West is significantly lower (0.212 opt Au versus 0.309 opt Au for the Range Front and 0.364 opt Au in the Ogee). Raw data for the CX-West zone shows most gold values within the grade shell to fall below 0.320 opt (3rd quartile). Figure 17-45 shows a histogram of the same values that have been transformed using natural log function. In Figure 17-45, the histogram shows a strong normal histogram curve, while the cumulative frequency curve shows a typical shape for a set of samples coming from a single population. Several potential outlier samples above 0.568 opt Au are present in the histogram of raw data. As outlined in Section 17.6, a decile analysis shows 28% of the gold contained in the range front grade shell occurs above the 90th percentile (0.486 opt Au), with 19% of the gold occurring in assays from the 80th percentile (0.348 opt Au). Visual inspection of the intercepts show a uniform and widespread distribution in all grade ranges, including those above .50 opt, with sufficient drilling surrounding the high grade assays to support and constrain the interpolation. The statistical analyses and visual observations indicate that no top cutting be required for the CX-West zone.

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Figure 17-44: CX-West Cummulative Probability

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Figure 17-45: CX-West Cummulative Frequency Curve

17.14.3 Block model parameters

Parameters for the block model covering the Range Front, Ogee, and CX-West grade shells are given in Table 17-16. The block model was created using 10-foot x 10-foot x 10-foot blocks with no rotation to the orientation of the model. Calculation of individual block grades within the model was done by separately constraining interpolations to the 0.1 opt Au grade shell solids for each zone. In other words, each zone was calculated separately using the grade shell as a constraint and using search parameters specific to the configuration and statistics for each zone.

For purposes of this resource estimate, separate search ellipses were created for each of the Ogee, Range Front, and CX-West zones. Search ellipse parameters were calculated for each grade shell by guiding the search along mineralizing controls and intersection zones as defined by geology. Ellipsoid parameters for each zone were generated by aligning the major and semi major axis parallel to the two longest dimensions of the mineralized domain (usually along strike and down dip of the controlling structures, with plunges oriented parallel to dominant plunges of intersection zones. Where two controlling structures are present, such as in the Ogee zone, the major and semi-major axis were aligned along the trend of the potential intersection trend. Distribution of mineralization down dip and along strike are generally equal in all zones such

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that the anisotropy ratios for the major and semi-major axis were set at 1:1, with the major/minor anisotropy set at 1:4. Table 17-17 show the search ellipse parameters used for each grade shell zone.

Table 17-16: Block Model Parameters

Parameter		Minimum	Maximum
Model size	Easting	8500	10850
	Northing	10350	12400
	Elevation	4000	5200
Block size		10’ x 10’ x 10’
Descritisation values		2 x 2 x 2
Rotation		Not rotated

Units for easting and northing are Pinson Local Mine grid coordinates.

Table 17-17: Search Ellipse Parameters

	Axial rotations				Anisotropy ratios
zone	bearing	plunge	dip	search radius	Major/Semi-major	Major/Minor
Upper RF	55	-67	-65	400	1	4
Ogee	50	-80	88	400	1	4
CX-West	68	0	84	400	1	4

Visual inspection of grade distribution within the block model was made on a section-by-section basis and on levels on a plan-by-level plan basis. Block grades were visually compared to drill assay values, both in terms of grade estimation and distribution, near and between drill holes. The visual inspection showed no irregularities in the block model estimations.

17.15 2007 Resource Estimates

Grade and ton estimates calculated for the Range Front, Ogee, and CX-West zones used a partial percentage value contained in the block model. The partial percentage is a value that reflects, on a volume basis, how much of the block is contained in the grade shell. The program calculates tons by multiplying expected tons for the block by this percentage to get the number of tons actually contained in the shell. For purposes of this resource estimate, the calculation was limited to blocks with at least 25% of the block volume contained in the shell. This would emulate a sub-block size of 5-foot x 5-foot x 5-foot for a 10-foot x 10-foot x 10-foot block.

In the Range Front zone, mineral resources were re-calculated for all drilling above 4,190 feet in elevation. The 4,190-foot elevation is the base of densest drilling, where drilling at 200-foot centers or less occurs above, and 400-foot centers or greater falls below this elevation. Additionally, Atna’s Phase 2 program did not drill below this level and therefore there is no new information to incorporate into a revised resource estimate below the 4,190-foot level. For the Ogee and CX-West zones, resource estimates were carried down to levels where similar drill densities generated the highest confidence in the geologic interpretation. In the CX-West zone estimates were cut at the 4,440-foot elevation, and in the Ogee zone at the 3,750-foot elevation. No drilling occurred in the CX resource during Atna’s Phase 2 program and the resource estimate for that zone remain unchanged from the March 2005 estimate.

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Resource classifications were made based on distance from drill holes (following the parameters used in the 2005 estimate), but were modified based on density of drilling and rock type. In zones where drill spacings were 200 feet or less, blocks were tagged in the model as measured resources if the nearest drill hole for estimation were less than 50 feet in actual distance from the block, and for indicated resources at less than or equal to 100 feet from the nearest drill hole. Inferred resources were calculated at distances greater than 100 feet. In zones where drill density was greater than 200 feet, all blocks were tagged as inferred resources, with no measured resources calculated. Indicated resources were calculated in zones where drill spacing was greater than 200 feet only if 3 or more holes were drilled on a grouping with separations less than 300 feet.

Rock type played a significant role in classifying resources in the upper Range Front. Particulary in the zone of intense shattering between the Range Front fault and the hanging wall splays that separate Upper Comus Shale from the Lower Comus Limestone. In this area, rock quality is deemed poor and may not provide stable conditions for mining. All resources hosted within the Upper Comus Shale along the Range Front zone have been classified as Inferred for purposes of the estimates in this technical report.

17.15.1 2007 Range Front Resource Estimate

Table 17-18 provides estimates on the classification, average gold grade, and tons for resources in the upper Range Front zone., and Figure 17-46 shows the distribution of grade within the zone.

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Table 17-18: 2007 Revised Range Front Resource Estimate

*Category*	*Cut-off* *(Au opt)*	*Tons* *(x 1,000)*	*Grade* *(Au opt)*	*Contain Au* *(x 1,000 oz)*
*Upper Range Front Zone (Surface to 4,200-foot level)*
Measured	0.20	278	.419	116.5
	0.25	228.9	0.461	105.5
	0.30	198.4	0.490	97.2
Indicated	0.20	307	0.391	120.04
	0.25	268	0.416	111.5
	0.30	230.2	0.438	100.8
Measured + Indicated	0.20	585.04	0.404	236.54
	0.25	496.9	0.436	216.6
	0.30	428.6	0.462	198

Inferred	0.20	242.14	0.313	75.8
	0.25	165.6	0.355	58.8
	0.30	110.38	0.397	43.8
*Lower Range Front Zone (below the 4200-foot level, March 2005 block model)*
Measured	0.2	0	0	0
	0.25	0	0	0
	0.3	0	0	0
Indicated	0.2	226.64	0.356	80.68
	0.25	205.4	0.368	75.6
	0.3	172.2	0.387	66.64
Measured + Indicated	0.2	226.64	0.356	80.68
	0.25	205.4	0.368	75.6
	0.3	172.2	0.387	66.64

Inferred	0.2	2,382.75	0.354	843.5
	0.25	1,914.5	0.375	727.94
	0.3	1,190.2	0.421	501.1

(Base case cut-off grade = 0.20opt Au)

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Figure 17-46: Gold grade distribution in the Upper Range Front zone.

Looking Southwest.

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17.15.2 2007 Ogee Zone Resource Estimate

Table 17-19 provides estimates on the classification, average gold grade, and tons for resources in the Ogee zone., and Figure 17-47 shows the distribution of grade within the zone looking northeast along the trend of the the CX-West fault zone. Figure 17-48 shows the Ogee zone grade distribution looking to the north along the trend of the Ogee fault zone.

Table 17-19: Ogee Zone Resource Summary

*Category*	*Cut-off* *(Au opt)*	*Tons* *(x 1,000)*	*Grade* *(Au opt)*	*Contained Au* *(1,000 x oz)*
Measured	0.20	438.5	0.632	277.1
	0.25	415.3	0.655	272
	0.30	387	0.682	263.9
Indicated	0.20	352.4	0.570	200.9
	0.25	331.7	0.592	196.4
	0.30	315.2	0.608	191.6
Measured + Indicated	0.20	790.8	0.604	477.6
	0.25	746.9	0.626	467.6
	0.30	702.2	0.649	455.7

Inferred	0.20	90.6	0.473	42.8
	0.25	80.7	0.503	40.6
	0.30	73.5	0.526	38.7

(Base case cut-off grade = 0.20opt Au)

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Figure 17-47: Gold grade distribution in the Ogee zone

(Looking northeast-along the CX-West fault zone trend)

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Figure 17-48: Gold grade distribution in the Ogee zone

(looking south, on strike of the Ogee fault trend)

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17.15.3 CX-West Resource Estimate

Table 17-20 provides estimates on the classification, average gold grade, and tons for resources in the Ogee zone and Figure 17-49 shows the distribution of grade within the zone.

Table 17-20: CX-West Resource Estimate

*Category*	*Cut-off* *(Au opt)*	*Tons* *(x 1,000)*	*Grade* *(Au opt)*	*Contained Au* *(x 1,000 oz)*
Measured	0.2	116.9	0.290	33.9
	0.25	79.8	0.321	25.6
	0.3	39.9	0.368	14.7
Indicated	0.2	40.5	0.270	10.9
	0.25	26.9	0.292	7.85
	0.3	11	0.321	3.5
Measured + Indicated	0.2	157.3	0.284	44.7
	0.25	106.7	0.314	33.5
	0.3	50.9	0.352	17.9

Inferred	0.2	1.07	0.229	.25
	0.25	.23	0.266	.06
	0.3	0	0	0

(Base case cut-off grade = 0.20opt Au)

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Figure 17-49: Grade distribution in CX-West Fault

(looking southwest)

17.15.4 Comparison of March 2005 Estimate with 2007 Estimate

The Phase 2 program was successful in converting 111,000 tons from inferred and indicated into the measured category within the Range Front zone resource. Significant tonnage in the inferred and indicated categories were eliminated in the upper Range Front zone as drilling showed the upper shale-hosted zones to be less continuous than previously interpreted. The vast majority of the measured and indicated resources added during the Phase 2 program were due to the discovery and delineation of the Ogee mineral resource and detailed drilling in the Range Front zones. The Ogee, with its higher grades, had the most significant impact on the overall resource grade. An additional 117,000 tons were added to the measured category with the addition of the CX-West resource. Table 17-21 summarizes the total project mineral resource. Measured plus Indicated resources, at a 0.20 oz/ton gold cut-off, total 2.505 million tons at an average grade of 0.424 oz/ton gold containing 1.063 million ounces of gold. Inferred resources, at a 0.20 oz/ton gold cut-off, total 3.374 million tons at an average grade of 0.340 oz/ton gold containing 1.146 million ounces of gold.

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Table 17-21: Combined Resource Summary (RF, CX, Ogee, CX-West)

*Category*	*Cut-off* *(Au opt)*	*Tons* *(x 1,000)*	*Grade* *(Au opt)*	*Contained Au* *(x 1,000 oz)*
Measured	0.20	1,152.4	0.454	523.2
	0.25	937	0.508	475.6
	0.30	755.3	0.565	426.5
Indicated	0.20	1,353.5	0.399	540.6
	0.25	1110	0.438	485.8
	0.30	884	0.480	525
Measured + Indicated	0.20	2,505	0.424	1,063
	0.25	2045	0.469	960
	0.30	1640	0.520	852.7

Inferred	0.20	3,374.5	0.340	1,146.6
	0.25	2,652	0.364	964.7
	0.30	1,507	0.419	631.7

(Base case cut-off grade = 0.20opt Au)

Comparison of the March 2005 and 2007 resource calculations above the 4200-foot elevation in the Range Front zone is shown in Table 17-22 for the 0.20 oz/ton gold cut-off. Within all categories, approximately 190,000 tons were eliminated from the block model within this zone. The reduction in tonnage is due to the re-interpretation of structure and the less pervasive nature of mineralization within the upper Comus shale package, which was only broadly drilled at the end of the Phase 1 effort. Grades however have increased almost 19% from the March 2005 estimate, going from average grade of 0.318 oz/ton gold to the new estimated grade of 0.378 oz/ton gold. The program was successful in moving 211,000 tons from the Inferred category into the Measured and Indicated categories, giving a total Measured and Indicated resource of 585,000 tons grading 0.404 oz/ton gold in the Range Front Zone..

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Table 17-22: March/2005 versus 2007 Resource-Upper Range Front

*Category*	*Cut-off* *(Au opt)*	*Tons* *(x 1,000)*	*Grade* *(Au opt)*	*Contain Au* *(x 1,000 oz)*
Measured-2005	0.20	0	0	0
Measured-2007	0.20	278	0.419	116.5
Percent change		+100%	n/a	+100%

Indicated-2005	0.20	374.4	0.343	128.2
Indicated-2007	0.20	307	0.391	120.4
Percent change		-18%	+14%	-6.1%

Inferred-2005	0.20	643.0	0.303	195.3
Inferred-2007	0.20	242.14	0.313	75.8
Percent change		-62.3%	+2.9%	-61.2%

Total-2005	0.20	1,017.4	0.318	323.5
Total-2007	0.20	827	0.378	312.6
Percent change		-18.7%	+18.9%	-3.4%

(Base case cut-off grade = 0.20opt Au)

Appendix 17 includes level plans of the Range Front, Ogee, and CX-West block models which display in greater detail the configuration and distribution of grades within the revised mineral resource estimate which is the subject of this technical report.

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18.0 Other Relevant Data

This section discusses the preliminary analysis of important issues related to the hypothetical development of the Pinson underground mineral resources. It is included to inform readers of this technical report about the type of mining and development that may occur at the project should all or part of the current mineral resource be successfully converted to proven and probable minable reserves which may be economically extracted. Readers are cautioned that there are currently no minable reserves at the Pinson Project and there is no assurance that Atna’s work on the project will ultimately result in the conversion of part or any of the current resources to a minable reserve. Additionally, readers are cautioned that the Pinson resources have not been the subject of an economic evaluation, pre-feasibility study or feasibility study and therefore do not currently have any demonstrated economic viability.

18.1 Potential Mining Methods

A mining method for any underground deposit is selected on the basis of geological modeling of the ore zones to determine width, height, dip angle and overall volume of the modeled ore zones. Other primary factors used in deciding which mining technique will be most effective are the geotechnical characteristics of the host and surrounding rock and general continuity of the ore blocks. A number of other factors including maximizing productivity, minimizing costs, optimizing of ore reserve recovery, minimizing dilution by waste rock and assuring a safe operating environment are also factored into the decision.

The preliminary analysis of the possible mining method to utilize at Pinson examined the existing Pinson ore bodies mined by open pit methods. This allowed analysis of hanging and footwall rock characteristics, behavior of major fault systems and the general dip of mineral bearing zones. Building on the knowledge of existing data from the two adjacent open pits and adding it to the extensive drilling program discussed in the prior sections has allowed a detailed geologic model to be developed that provides the key elements necessary to select a mining method for the Pinson underground resource, should it be converted into a minable reserve after completion of a feasibility study.

Several of the factors that are defining the Pinson mining method include:

·	It is expected that the mineral zones will generally dip at an angle of 45 to 60 degrees.

·	The ground conditions in the mineral zone are sufficiently broken as to not support large openings.

·	The block modeling is identifying relatively discontinuous pods of high-grade mineralization.

These factors indicate that the mining method must be restricted to a very selective mining method. The preliminary choice mining to employ at Pinson appears to be underhand cut and fill which is the method commonly used at the Getchell underground mine and throughout the Carlin trend in similar geologic environments.

18.2 Preliminary Geotechnical Evaluations

A preliminary geotechnical evaluation was performed by Mr. J. Roland Tosney of Minefill Services, Inc. during November and December of 2005. The purpose of the evaluation was to assess the Pinson rock mass quality in areas where waste development or ore mining may occur. Potential mining at Pinson is likely to occur primarily within the Ordovician Comus formation. The Lower Comus, Upper Comus and mineralized portions of the Comus formation form the preliminary basis for differentiating between the various geotechnical units at Pinson.

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As anticipated, the Range Front and CX zone are characterized by “very poor” to “extremely poor” rock mass characteristics (RMR < 30, Q < 0.2) consistent with the fault related deposit geology. As a result, it is very likely that underhand cut-and-fill will be the most suitable mining method for Pinson. Based on this information, drift dimensions in the mineralized zone will have a limited span on the order of 10-12 feet in width and height. It is likely that 2 to 3.5 inches of fiber reinforced shotcrete along with patterned rock bolting on 4-foot centers will be necessary for ground support within open stopes.

Note that drift dimensions will increase and ground support requirements will decrease for waste development work occurring away from fault/mineralized zones, within the more competent areas of the Upper and Lower Comus rock units. Patterned bolting and screening, consistent with current practices, will likely be sufficient in these areas.

For secondary ore cuts (i.e. cuts excavated beneath cemented rockfill) stope dimensions are determined based on the required strength of the cemented rockfill. Mining widths under the cemented rockfill are anticipated to be a maximum of15 feet (See backfill section below).

18.3 Potential Drift and Fill Mining Method Layout

Layout of the drift and fill stopes varies dependent upon the shape of the mineralized body to be exploited. In general the panels are designed for maximum length. This usually means drifting in the strike direction of the mineralized zone. The maximum length is considered to be in the 400 foot range at most operations in Nevada in similar environments. Mining in this manner is the most efficient because downtime is decreased between panel completion, backfilling and final surveys. Ideally the first drift is driven along the footwall contact of the mineralized zone. Once the footwall drift is completed and backfilled, a new panel is started toward the hanging wall alongside the footwall drift. Additional parallel drifting will follow until the ore has been extracted up to the hanging wall contact (See Figure 18-1 and 18-2 for examples).

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Figure 18-1: Schematic 4500-Level – Range Front and Ogee Zones

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Figure 18-2: Schematic 4700 Level Production Plan – Range Front and Ogee Zones

It is expected that some areas of the hypothetical mine the mineralized zone will have a strike length roughly the same as the length in the dip direction. For this type of drift and fill stoping a different layout is used. The initial panels are driven on either the footwall or hanging wall in mineralized zones. Cross cut panels are then driven on 60^° turns from the main access. This angle allows for effective access of equipment but also minimizes potential "nose" pillars. Mining and backfilling is always done on "retreat", from the end of the ore zone out toward the access.

Upon completion of mining and backfilling in an entire work area, the area will be re-accessed by declining in the production access drift. The back of the newly developed area is the same elevation as the sill of the prior cut, therefore, the back will consist of the compacted, cemented rockfill. This sequence will be repeated to allow for six separate ore accesses to be developed from one main development drift. From the main waste declines, working elevations will be established every 90 feet which allows for the six 15-foot high production cuts (See Figure 18-3).

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Figure 18-3: Schematic Production Access Development

A second method may be utilized where sufficiently continuous ore zones when dips are greater than 45 degrees. In this case longhole stoping and development will be utilized providing improved productivity. While this method provides for more efficient mining it is expected that only 10% to 20% of the mineralized zones would support this method due to wall rock conditions adjacent to the mineral zones. The Ogee Zone, in particular, may have the geometry and rock conditions conducive to longhole stoping. The CX zone is also being considered for this type of stoping.

18.4 Description of Potential Mine Development

There are currently four areas that will require access for mining of the Pinson underground mineral resources (if they are successfully converted to mineral reserves), the CX zone, the Range Front zone to the northwest of the CX zone, and the CX-West and the Ogee zones situated between the Range Front and CX mineralized zones. The zones will be accessed through a series of drifts and declines that start from the 4765 bench of the CX open pit mine (the current exploration drift access portal).

Hypothetically, production ore would be extracted from the stopes using 4-6 yard LHD equipment. Both ore and waste material will be transported in 15 to 25 ton trucks approximately 2,000 to 10,000 feet (dependent upon stope location and elevation) up a +14% incline to the surface. The same trucks will backhaul engineered backfill for the drift & fill stopes that are being backfilled.

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The primary production drift for the Range Front mineral zone has been partially developed as part of the Phase 2 exploration project. Additional capital development will be required to develop the required production haulage access drifts, access to the identified stoping areas, secondary escape-way, and additional decline development to access new mine levels. In addition, required support facilities including maintenance shops, substations, lunch rooms, muck bays, would also be necessary prior to production commencing.

Development of hypothetical mine plans and operating conditions suggest that Pinson may be able to attain a production rate averaging 170 tons per day per stope/active working face. This production rate includes time for backfilling as well as downtime over the life of the stope. Short-term rates per stope can be as high as 350 tons per day. For future economic evaluation purposes, the Pinson project should assume that the underground mine would actively work eight drift and fill stopes at all times, 6 to 7 in production and 1 to 2 backfilling. This will provide a total production of approximately 800-1000 tons per day. These rates are comparable to other mines in the area and appear at this point to be potentially achievable given the location and current understanding of the mineral resources.

18.5 Analysis of Ground Support Requirements

Experience gained during the Phase 2 program drifting indicate that the initial drift and fill panels will be supported with 6-foot split set bolts and 6-foot x 9-foot wire panels with 4 inch squares. Ribs are bolted down to within 7 feet of the sill. With the exception of the primary cuts in ore, shotcrete is applied if necessary but is not part of the normal cycle for drift and fill mining. Once the initial stope is completed and mining starts beneath the backfill, very little mechanical ground support should be required. When mucking the drift and fill rounds, the loader operator will scrape the back with the bucket to ensure that there are no loose pieces of backfill overhead.

18.6 Required Backfill & Equipment

Initial cement rock backfill requirements have been assessed by Mr. J. Roland Tosney of Minefill Services Inc. (referenced in References section of this report). Since the initial cuts in ore are envisioned to have moderate-narrow spans, the initial recommendation is to produce backfill meeting an approximate specification of 500 psi compressive strength. Initial testing of the exposed upper Comus shale unit in the CX-West open pit also indicate this area to be an acceptable source of aggregate for the cemented rockfill.

The backfill will be a rock/cement mixture with cement representing 4-5.5% of the weight. The aggregate will be crushed to approximately -3 inch by an in-pit crushing and screening facility then hauled to a surge pile at the CX Pit floor near the portal entrance. The aggregate will then be fed by loader to a batch plant located on the 4760 bench of the CX pit where a measured mixture of water and cement will be batched with the aggregate. Addition of flyash as a substitute for cement is also under consideration. The backfill is required to produce a compressive strength between 300 - 500 psi to maintain a safe work environment.

The backfill will be mixed near the portal and loaded into trucks for hauling into the mine. The haulage trucks that haul ore or waste out of the mine are then loaded with the backfill material for the return trip underground, maximizing the efficiency of the haulage equipment. The backfill material is discharged into the stope to be backfilled and then jammed into position using a LHD loader with a push plate attached to the front end.

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18.7 Dewatering

A dewatering program designed to lower the cone of influence of the existing water table will be conducted prior to mining mineralized zones. Test dewatering was part of the Phase 2 program and was being accomplished with a 200 hp submersible pump via a dewatering well drilled 640 feet deep in the bottom of the CX Pit. Future plans for deeper development (beyond the 4100-foot level) would require the deepening of the dewatering well to 1200 feet and installing a booster pump at the collar of the well. An additional back-up well is also being considered which would be drilled and in production prior to deepening the existing well. The current dewatering system averages 1200 gpm capacity and a groundwater drawdown rate of about 20 feet per month was recorded during the test dewatering program in Phase 2. Water discharge is via a 12 inch HDPE pipe which transports the water from the well head up a 400 foot vertical rise out of the CX pit then approximately 12,000 feet to an approved temporary discharge point near the Granite Creek drainage channel. In the future, the fully permitted rapid infiltration basins will be constructed in the Granite Creek alluvial basin and the discharge will be relocated accordingly.

Normal mine water handling will be in the 400 to 600 gpm range, while short-term water handling could reach 2,500 gpm. As the water pockets are localized and can be controlled to a certain degree by grouting and shotcreting, the dewatering plan will be based on a maximum of 400 gpm. An in-pit settling sump will be constructed to collect, clarify, and then pump the mine water to the approved infiltration basin.

18.9 Infrastructure

18.9.1 Water Supply

Two existing fresh water wells, located east of the Getchell Road are utilized to supply potable water to the mine site via a buried steel pipe system. The wells are equipped with 200 HP line shaft pumps which are operated with the use of diesel generator sets.

Additional water supply for fire suppression, dust control, and other onsite activities is contained in a 175,000 gallon water tank, the source of which is supplied by the two fresh water wells. In the future, APW-1 dewatering well has the potential to become the feed source for all non-potable needs at the mine site in addition to its primary function as a dewatering well.

18.9.2 Power Supply

A new electrical distribution system was installed in the summer of 2005. Line electrical power is supplied by Sierra Pacific Power via the existing 120 Kva line that runs northerly along the Getchell Road corridor through the Pinson property. From the Getchell Road, power is suspended on power poles approximately 1500 feet to the Pinson substation where the power is stepped down to 13.8 Kva. From the Pinson substation, the 13.8 Kva main power line is routed approximately 5000 feet along power poles to the edge then into the bottom of the CX Pit to an in-pit substation. From this substation, electrical power is distributed to the underground mine and any mine related facilities on surface requiring power.

All potential surface facilities will utilize the same 13.8 Kva line feed for their respective power needs in the future.

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18.9.3 Buildings

Construction of a new office/dry house/core logging facility commenced in the summer of 2005 and was completed in mid-2006. The building is 125 feet by 75 feet or approximately 9300 square feet. This building will house the administrative, operations, and geology staffs as well as function as a shower and change area and training area for the mine crews.

The only two other surface facilities currently under consideration are an onsite lab/sample preparation building and a truck weighing facility to track ore shipments to a third-party toll milling facility. The need for rapid assay turn-a-round when the mine is in production may dictate the need to have an onsite assay lab, but this determination has not been made to date. Obviously, the scale-truck weighting facility will be required if the project is taken through feasibility and into commercial production, but is not currently required at this stage of the project.

18.10 Environmental and Socio-Economic

18.10.1 Environmental

Pinson Mining Company operated the Pinson Mine from 1980 through cessation of surface mining operations January 28, 1999. Components of Pinson Mine operations 1980-1999 included ten open pits as well as waste dumps, leach pads, tailings and plant facilities. Figure 18-4 is a map showing the location of the pits and other mine facilities. Pinson Mining Company pursued final permanent closure and reclamation of its components during 1999-2004 and much of the affected areas are at the point of final reclamation bond release. Post-closure monitoring of one reclaimed tailings impoundment, the heap leach pad fluid management system, site groundwater and lakes that formed in the CX and Mag open pits is ongoing and will continue as required by Nevada State regulatory authorities for an extended period and will result in an on-going cost to the project of approximately US$10,000 per year. Only minor other tasks remain to be completed. Outstanding tasks include completion of administration facility reclamation by burying the administration and reagent storage building foundations, recontouring the administration facility area and performing final revegetation of this area.

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Figure 18-4: Mine Facilities

Mining operations in the Mag pit and the CX pit intercepted the existing water table and mine dewatering operations were conducted to lower the water table below the lowest working level of the open pits. Following cessation of surface mining operations and dewatering operations, permanent lakes formed in the bottoms of the two open pits.

The Mag pit lake has approximately 250 million gallons of water impounded and a stable final surface water elevation, subject to seasonal fluctuations resulting from precipitation onto the surface and evaporation from the surface of approximately 4671 ft amsl. The water in the upper 90 feet of the 130-foot deep lake meets Nevada MCL for drinking water.

Approximately 11 million gallons were impounded in the CX pit lake to a surface elevation of 4714 ft amsl by December, 2005. The water impounded in the 28-foot deep lake met Nevada MCL for drinking water.

Atna began dewatering the CX pit lake and the surrounding Pinson Fault Shear Zone hydrologic unit December 13, 2005. The dewatering operations are intended to lower the water level in the shear zone to facilitate underground exploration and mining operations to at least an elevation of 4200 ft amsl. As of January 1, 2006, dewatering operations were ongoing at a rate of 1150

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gpm and the water elevation in the shear zone had lowered 15 ft to 4699 ft amsl over the initial 18 days of dewatering operations. The CX pit lake was entirely eliminated by the end of January 2006 by virtue of the water level in the Shear Zone being drawn below the 4686 ft amsl floor elevation of the CX open pit. Cessation of dewatering test work has now allowed the recovery of the pit lake to approximately 4710 amsl.

18.10.2 Environmental Permitting

Various Federal, State of Nevada and Humboldt County permits are required for authorization to conduct surface and underground exploration and mining operations at the Pinson Underground Project. The following table identifies those key authorizations and the current status of authorizations. It should be noted that no new disturbance on Federal lands at the Pinson mine is proposed under current operating plans.

Table 18-1: Key Permits

TYPE OF PERMIT	STATUS	COMMENTS
Water Pollution Control Permit Pinson Exploration Project	In-place	Modification required for ore production increase.
Water Pollution Control Permit Pinson Infiltration Project	In-place	Authorizes discharge of mine dewatering water to Rapid Infiltration Basins (RIBs).
Water Pollution Control Permit Pinson Post-Closure	In-place	Authorizes post-closure monitoring of previous Pinson mine closed components.
Reclamation Permit for Pinson Mine Exploration Project	In-place	Modification required for underground backfill source disturbance.
Reclamation Permit for Pinson Mine (prior disturbance)	In-place	Authorizes reclamation of disturbances remaining from Pinson Mine operations 1980-2004.
Mining Plan of Operations	In-place	Authorizes operations and reclamation of disturbances 1980-2004 on Federal lands at the Pinson mine.
Class III Air Quality Operating Permit	In-place	Authorizes limited operation of backfill and shotcrete plants, surface area disturbance and genset operation.
Surface Area Disturbance Air Quality Operating Permit	In-place	Authorizes emissions from surface area disturbances remaining from Pinson mine operations 1980-2004.
Class II Air Quality Operating Permit	In-place	Class II AQOP required for increased emissions from backfill and shotcrete plants, crusher, surface area disturbance.
Temporary Authorization to Discharge	In-place	Expires 4/13/06.
Notice-level authorization for monitoring wells on BLM land	In-place	Authorizes construction and use of groundwater monitoring wells downgradient of RIBs.
Mining Storm water General Discharge Permit	Application 2Q06	New general permit covering Atna operations.
Mining Stormwater General Discharge Permit (Pinson mine prior operations)	In-place	Authorizes stormwater discharges from remaining Pinson mine operations 1980-2004.

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Table 18-2: Other Required Permits for Mining

TYPE OF PERMIT	STATUS	COMMENTS
USEPA Hazardous Waste Generator ID Number	Need to be determined	Required if facility generates hazardous waste. May continue to utilize current Pinson Mining Company EPA ID#.
Toxic Release Inventory Facility Identification #	Need to be determined	May not be required if no onsite gold production.
Water Appropriation Permits	In-place	Modification to appropriations may be necessary depending on future dewatering needs. Current appropriations adequate for current and near-term projected needs.
Hazardous Materials Storage Permit	Need to be determined	Authorizes onsite storage of hazardous materials.
Public Water Supply Permit	Need to be determined	Authorizes provision of public water supply if onsite man-hours exceeds 52,000 in a one-year period.
Small Facility Septic System Permit	In-place	Modification approved for hookup of new building to existing system.

18.10.3 Population, Demographics & Ethnicity

The Humboldt County population base is about 20,000. The most densely populated portion of the county is the town of Winnemucca (pop. 7,500). The median age of Humboldt County residents is significantly younger than most other Nevada counties (early 30s vs. early 40s). The ethnicity of Humboldt County is predominantly Caucasian, and projected to remain that way.

18.10.4 Employment

Employment surveys by the State of Nevada in May, 2004 indicate total employment in the county is just under 7,000 persons (Nevada Dept of Employment, Training & Rehabilitation, 2004). The employment base in Humboldt County is nearly distributed according to the "80-20" rule: nearly 80 % of the employment is accounted for by 20 % of the employers (Sierra Pacific Economic Development Office, 2004). There are 5 major employers employing over 250 workers each (the school district and several mines), 13 employing over 100 each, and more than 100 other small businesses. Much of the small-business employment is tourist- and gaming-service oriented.

18.10.5 Workforce Qualifications

The region around Winnemucca has a long history of mining activity. Several world-class or long-lived gold mines operate in a 50-mile radius from Winnemucca, and this fact alone has created a large body of skilled and experienced manpower. A number of contractors are available locally to meet construction and repair needs, supply industrial equipment and parts, and provide onsite technical services including exploration, road construction, excavation, drilling, contract mining, mill operation, and remediation and closure. Those services not immediately available are obtainable from Reno, 175 miles to the west, or from Elko, 150 miles to the east.

190

19.0 Conclusions

Mineralization located within the CX, CX-West, Ogee and Range Front mineralized zones at the Pinson Mine project represent significant bodies of gold-mineralized rocks with characteristics similar to the economic and productive Carlin-type gold systems currently in production in underground mines located in the Getchell Gold Belt and other districts in Northern Nevada.

Gold mineralization is hosted by the same stratigraphic units hosting the Getchell deposit to the north of Pinson. Pinson’s mineralization has a similar structural control to that present at the Getchell Mine, five miles to the north, occurring within the network of faults making up the horst bounding, Basin and Range fault zone along the southeastern margin of the Osgood Mountains. Gold grades within the CX, CX-West, Ogee, and Range Front zones are similar to grades at other underground mine properties in the region currently in production and the zones remain open in several directions.

Owing to the vast amount of information existing prior to Atna’s commencement of work at the property and data collected from Atna’s 88 surface and 48 underground drill holes (136 holes total), the geologic understanding of the mineral system’s configuration, structure and stratigraphic control are adequate to support the current resource model contained within this updated technical report.

Drilling is rather wide spread, in the deeper portions of the four gold mineralized zones that are the focus of Atna’s work (CX, CX-West, Range Front, and Ogee). Additional drilling recommended in conjunction with underground development and feasibility work will provide the additional data density required to bring additional portions of the inferred mineral resource into the measured and indicated categories and potentially into mineral reserves, once a mine plan and feasibility study are completed.

Atna’s re-assaying program on the existing pulps from earlier operator’s drilling confirmed the extent and analytical values within industry error standards. Atna’s sampling procedures of both core and reverse circulation drill holes have been conducted according to accepted industry standards. Logging procedures for both core and cuttings utilize similar approaches and techniques in use by all major mining companies and in the adjacent mine properties. Sample preparation and analysis by Inspectorate America of Reno, Nevada was performed using industry-standard fire assay methods analytical procedures, incorporating blind internal check samples, analytical standards, pulp replicates and blanks to insure reliability and reproducibility.

Atna has a written Quality Assurance and Quality Control procedure in place which includes the insertion of blind certified analytical standards, blank samples, duplicate samples, and replicate assays. The program includes the routine submission of the mineralized pulps to a second laboratory, ALS Chemex, of Reno, Nevada, for replicate analysis. Atna has taken and continues to take adequate quality control and assurance steps to insure the quality of the analytical data on the Pinson Property.

The analytical database utilized to produce the resource model contained within this Technical Report was audited with minor error rates.

Methods utilized to determine the March 2005 resources models for the Range Front and CX mineral zones at the Pinson Project by Robert Sim are consistent with practices in standard use in the industry. The revised resource models and estimates contained within this technical

191

report for the Range Front, CX-West and Ogee zones are consistent with the original work completed by Robert Sim and follow standards and methods currently utilized in the mining industry.

20.0 Recommendations

The current mineral resource base at Pinson is sufficient to warrant the completion of an economic evaluation, pre-feasibility or feasibility study of the viability of mining these gold resources. It is recommended that a feasibility study be completed on the currently defined measured and indicated resources with concurrent delineation drilling within the indicated portion of the resource to add additional resources that may also be economically viable and therefore converted to minable reserves during work on the feasibility study. The vast majority of this development drilling will require additional underground platforms to be developed and it is recommended that this work be completed contemporaneously with the feasibility study of the existing mineral resource.

The next phase of development should include the completion of mineral zone access drifts to the Range Front zone on the 4700 level and to the Ogee zone on the 4800 and 4700 levels. A centralized decline system, located between the two zones, should be advanced to provide mineral zone access below the 4800 level and will enable the establishment of diamond drill stations to delineate deeper portions of the both the Ogee and Range Front zones beyond the current measured and indicated resource. Upon completion of the mineral zone accesses to the Range Front and Ogee zones, work should include test mining to validate the mining method and ground conditions before adopting the proposed mining method in the feasibility study.

In order to execute the recommendations, the groundwater dewatering system will need to be completed to allow the groundwater table to be pumped below the mine workings ahead of the development of the decline and diamond drill stations. An in-pit mine water settling sump and a surface infiltration basin (both are currently permitted with regulator authorities) will need to be finished as soon as practical.

If this proposal is adopted, any required permit modifications should be processed as soon as practical.

192

References

Arehart, G.B., Chakurian, A.M., Tretbar, D.R., Christensen, J.N., McInnes, B.A. and Donelick, R.A., 2003, Evaluation of radioisotope dating of Carlin-type deposits in the Great Basin, western North America, and implications for deposit genesis: Economic Geology, v. 98, p. 235-248.

Benchmark Maps, 2003, Nevada road and recreation atlas: Medford, OR; 95 p.

Bristol, W., 2005, 2006, Internal company memoranda.

Bureau of Land Management (BLM, Winnemucca Field Office), 2001, Granite creek administrative draft environmental assessment N63-EA01-xx, prepared for Pinson Mining Company, 47 p.

Chevillon, V., Berentsen, E., Gingrich, M., Howald, B. and Zbinden, E., 2000, Geologic overview of the Getchell Gold Mine geology, exploration and ore deposits, Humboldt County, Nevada, in Crafford, A.E.J. (ed.). Geology and ore deposits 2000: the Great Basin and beyond: Geological Society of Nevada Symposium Proceedings, Field Trip # 9, Geology and ore deposits of the Getchell region, Humboldt County, Nevada, p. 113-121.

Cline, J.S., 2004, Introduction to Carlin-type deposits: Soc. Econ. Geologists Newsletter, No. 59 (Oct. 2004), p. 1.

Crafford, E.J., 2000, Overview of regional geology and tectonic setting of the Osgood Mountains region, Humboldt County, Nevada, in Crafford, A.E.J. (ed.). Geology and ore deposits 2000: the Great Basin and beyond: Geological Society of Nevada Symposium Proceedings, Field Trip # 9, Geology and ore deposits of the Getchell region, Humboldt County, Nevada, p. 69-83.

Davis, B.M., 1997, Some Methods of Producing Interval Estimates for Global and Local Resources: SME Preprint 97-5,4p.

Davis, D.A. and Tingley, J.V., 1999, Gold and silver resources in Nevada: Nev. Bureau of Mines & Geology Map 120.

Dawson Metallurgical Laboratories, Perry Allen-Metallurgist, 2005 & 2006, Memoranda regarding metallurgical test results and analysis, 2005 & 2005, Unpublished report to company.

Edmondo, G., 2004, 2005, 2006, Internal company memoranda.

Erickson, R.L. and Marsh, S.P., 1974, Geologic Map of the Iron Point Quadrangle, Humboldt County, NV: U.S. Geological Survey Geologic Quadrangle Map GQ-1175, scale 1:24,000.

Foster, J.M. and Kretschmer, E.L., 1991, Geology of the MAG deposit, Pinson Mine, Humboldt County, Nevada, in Raines, G.L., Lisle, R.E., Schafer, R.W., and Wilkinson, W.H. (eds.), Geology and ore deposits of the Great Basin: Geological Society of Nevada, 1990 Symposium Proceedings, p. 845-856.

Foster, J.M., 1994, Gold in arsenian framboidal pyrite in deep CX core hole DDH-1541: Unpublished Pinson Mining Company Report, June 21, 1994.

193

Hays, B., Koehler, S., Hart, K. and Whittle, T., 2004, Cortez Joint Venture Field Trip, in Carraher, R., Zbinden, E. and O'Malley, P. (eds.), Geological Society of Nevada Special Publication No. 40: New discoveries, exploration and mining activities along the central and southern Battle Mountain-Eureka trend, Lander and Eureka Counties, Nevada, p. 91-115.

Hofstra, A.H. and Cline, J.S., 2000, Characteristics and models for Carlin-type gold deposits: Reviews in Economic Geology, v. 13, p. 163-220.

Ilchik, R.P. and Barton, M.D., 1997, An amagmatic origin of Carlin-type gold deposits: Economic Geology, v. 92, p. 269-288.

Johnston, M.K. and Bessel, M.W., 2004, Carlin-type and distal-disseminated Au-Ag deposits: Related distal expressions of Eocene intrusive centers in north-central Nevada: Soc. Econ. Geologists Newsletter, No. 59 (Oct. 2004), p. 12-14.

Jones, A.E., 1991, Tectonic significance of Paleozoic and early Mesozoic terrane accretion in northern Nevada: University of California, Berkeley, Ph.D. dissertation, 256 p., 3 maps.

Kretschmer, E.L., circa 1983, Guidebook geology of the Pinson Mine. Unidentified report, 10 p.

MacKerrow, D.G., Whitney, M. and Ridgley, V., 1997, Geology at the Twin Creeks Mine, Humboldt County, Nevada: History, Development, Operations and Potential. Unpublished report.

Minefill Services Inc., J. Roland Tosney, November 2005, Atna Pinson Project – Review of Geotechnical Data and Preliminary Mining and Ground Support Recommendations. Unpublished technical memorandum.

Minefill Services Inc., J. Roland Tosney, December 2005, Atna Pinson Project – Cemented Rockfill Aggregate Assessment. Unpublished technical memorandum.

Madden-McGuire, D.J. and Marsh, S. P., 1991, Lower Paleozoic host rocks in the Getchell gold belt: Several distinct allocthons or a sequence of continuous sedimentation?: Geology, v.19, p. 489-492.

McLachlan, C.D., Struhsacker, E.M. and W.F.Thompson, 2000, The gold deposits of Pinson Mining Company: a review of the geology and mining history through 1999, Humboldt County, Nevada, in Crafford, A.E.J. (ed.). Geology and ore deposits 2000: the Great Basin and beyond: Geological Society of Nevada Symposium Proceedings, Field Trip # 9, Geology and ore deposits of the Getchell region, Humboldt County, Nevada, p. 123-153.

Nevada Department of Employment, Training & Rehabilitation, 2004. (http://detr.state.nv.us/lmi/data/wages/TOC011.htm)

Ridgley, V., 2004, 2005, 2006, Internal company memoranda.

Seedorff, E. and Barton, M.D., 2004, Enigmatic origin of Carlin-type deposits: an amagmatic solution?: Soc. Econ. Geologists Newsletter, No. 59 (Oct. 2004), p. 14-16.

Sim, Robert, March 2005, Technical Report on the Pinson Gold Property, Humboldt County, Nevada, Atna Resources internal report (also filed on SEDAR).

194

Sim, Robert, December 2005, Revised Technical Report on the Pinson Gold Property, Humboldt County, Nevada, Atna Resources internal report (also filed on SEDAR).

Sierra Pacific Economic Development Office, 2004. (http://econdev.sierrapacific.com/sppc/county/humboldt/bus_overview/)

Stenger, D.P., Kesler. S.E., Peltonen, D.R. and Tapper, C.J., 1998, Deposition of gold in Carlin-type deposits: The role of sulfidation and decarbonation at Twin Creeks, Nevada: Economic Geology, v. 93, p. 201-215.

Stewart, J.H., 1980, Geology of Nevada, Nevada Bureau of Mines and Geology Special Publication 4, 136 p.

Stoddard, S.W., Feeney, S., Verre, K., Wong, L. and Harris, T.R., 1999, Humboldt County Economic Development Profile: University of Nevada, Reno, University Center for Economic Development, Technical Report 99/2000-11, 59 p. (http://www.ag.unr.edu/uced/reports/technicalreports/fy1999_2000/992000_11rpt.pdf)

Thompson, W.F., 2003, [Geologic] Scoping Report for the Pinson Properties: Internal company report (March 2, 2003)

Tingley, J.V., 1998, Mining districts of Nevada: Nev. Bureau of Mines & Geology Report 47, 128 p.

Wallace, A., and Wittkopp, R.W., 1983, The mode of occurrence of gold at the Pinson Mine as determined by microprobe analyses, and metallurgical implications: Unpublished Pinson Mining Company Report.

Water Management Consultants, Inc., 1998, Results of Hydrogeologic and geochemical studies at Pinson and provisional closure plan for the CX and MAG pits: Consulting report, July 1998.

Winnemucca, NV web site (http://www.winnemucca.nv.us/)

195

Description of Activity	Cost (US$)
Underground Development Drifting and Test Mining (2,000 feet)	$2,500,000
Underground Diamond Drilling (20,000 feet)	$1,000,000
In-Pit Mine Water Sump Construction	$500,000
Rapid Infiltration Basin Construction	$300,000
Feasibility Study Contractor(s)	$200,000
Metallurgical Studies	$100,000
Engineering Studies and underground mine design	$100,000
Permitting	$100,000
Project Personnel	$300,000
Dewatering and project power costs	$100,000
Miscellaneous and Contingency	$300,000
Estimated Proposed Phase 3 cost	$5,500,000

Parcel Location	Property Description	Acreage	Royalty Description
Township 38 North, Range 41 East
Section 36	18-lode claims-PMC	360	5.0% NSR; and $0.50 per ton of ore produced.
Township 38 North, Range 42 East
Section 16	12-lode claims	264	5.5% NSR
Section 16	3-lode claims	56	5.0% NSR
Section 20	28-lode claims	460	No Royalty Burden
	5-lode claims	102	4 -5% NSR on precious metals, 2.5% NSR on base metals, & $1.00 to $1.50/ton for barite or barium based on specific gravity of ore.
	4-lode claims	78	5.5% NSR
Section 21	640 acres of private land	640	5.5% NSR
Section 22	18-lode claims	360	5.0% NSR
Section 23	640 acres of private land	640	7.5% NSR
Section 24	36-lode claims	640	2.5% NSR 1.0% NSR capped at $135,000
Section 25	480 acres of private land	480	7.5% NSR
Section 26	36-lode claims	640	2.5% NSR) 1.0% NSR capped at $135,000
Section 27	640 acres of private land	640	7.5% NSR
Section 28 (see footnote 1)	23-lode claims	400	5.0% NSR
	7-lode claims	120	5.5% NSR
	120 acres, private land	50 (net)	2.0% NSR
	120 acres, private land	70 (net)	5.0% NSR
Section 29 (see footnote 1)	640 acres of private land	640	5.5% NSR
Section 30	32-lode claims- PMC	545	No Royalty burden
Section 30	5-lode claims-PMC	95	5.0% NSR
Section 32 (see footnote 1)	18-lode claims	370	2.0% NSR
Section 32 (see footnote 1)	20-lode claims	270	5.0% NSR
Section 33 (see footnote 1)	640 acres of private land	640	7.5% NSR
Section 34	36-lode claims	640	2.5% NSR 1.0% NSR capped at $135,000
Section 36	18-lode claims	320	2.5% NSR 1.0% NSR-capped at $135,000
Township 38 North, Range 43 East
Section 30	36-lode claims	640	2.5% NSR 1.0% NSR-capped at $135,000
Township 37 North, Range 41 East
Section 12	25-lode claims	450	5.0% NSR $0.50 per ton of ore produced
Section 13	40-acre fee land	40	5.0% - 5.5% NSR
Section 13	320-acre fee land	320	5.0% - 5.5% NSR
Section 14	24-lode claims	438	5.0% NSR and $0.50 per ton of ore produced
Township 37 North, Range 42 East
Section 4	10-lode claims	200	No Royalty burden
	20-lode claims	320	5.0% NSR
	8-lode claims	120	5.0% NSR
Section 6	36-lode claims	640	5.0% NSR
Section 8	30-lode claims	520	5.0% NSR
Section 8	6-lode claims	120	No Royalty
Section 16	3-lode claims	50	5.0% NSR
Section 18	18-lode claims	360	5.0% NSR and $0.50 per ton of ore produced
Section 18	18-lode claims	280	2.5% NSR 1.0% NSR-capped at $135,000
	Total Acreage	14,018

Month	Extreme Low	Average Low	Average High	Extreme High
January	– 1 F (– 18 C)	17 F (– 8 C)	41 F ( 5 C)	57 F (14 C)
April	16 F (– 9 C)	31 F (– 1 C)	62 F (17 C)	80 F (27 C)
July	38 F (3 C)	53 F (12 C)	91 F (33 C)	102 F (39 C)
October	15 F (– 9 C)	31 F (– 1 C)	67 F (20 C)	84 F (29 C)

Deposit	Year of Discovery		Years in Production	Initial Reserves			Gold Produced (troy oz)		References
Deposit	Year of Discovery		Years in Production	Short Tons	Oz/Ton Au	Contained Ounces Au	Mill Ore	Leach Ore	References

Gold deposits of the Pinson Mining Company (PMC)
A	1963, 1971		1980 - 1985	2,500,000	0.108	270,000	369,753	83,469	Hill, 1971, PMC, 1993
B	1971		1982 - 1988	3,400,000	0.050	170,000	included	above	as above
C	1982		1988 - 1996	233,000	0.017	3,961	10,773	na	PMC, 1993, 1999
CX	1982		1990 - 1999	1,684,000	0.070	117,880	83,951	33,884	PMC, 1993, 1999
CX-West	1993		1994 - 1999			0	3,962	in CX	PMC, 1996, 1999
Mag (mill ore)	1984		1987 - 1999	4,300,000	0.080	344,000	301,255	na	PMC, 199 , 1999
Mag (leach ore)				2,300,000	0.030	69,000	na	59,741	Foster and Kretschmer, 1991, PMC, 1999
Felix	1972		1989 - 1992	355,000	0.030	10,650	1,133	11,641	PMC, 1993, 1999
Blue Bell	1972, 1983		1993 - 1994	228,000	0.072	16,416	17,014	1,085	PMC, 1993, 1999
Pacific	1984		1992 - 1993	130,000	0.048	6,240	4,939	2,607	PMC, 1993, 1999
Pinson Mine			08/1999 - 12/1999				0	2,141	PMC, 1999
Pinson Mine combined production						1,008,147	792,780	194,568	Total Pinson Mine Production 987,348 ounces gold


Prior gold production on PMC properties
Ogee and Pinson		1945	1949 - 1950					~10,000	Hill, 1971

OGEE ZONE CHANNEL SAMPLE ASSAYS
Sample No.	From(feet)	To(feet)	Length(feet)	Gold oz/ton(gram/tonne Au)
NORTH RIB
RFUG-055	76	81	5	0.144 (4.94)
RFUG-056	81	85	4	0.445 (15.26)
RFUG-059	85	88	3	0.274 (9.39)
RFUG-061	88	93	5	1.448 (49.65)
RFUG-063	93	97	4	0.176 (6.03)
RFUG-064	97	101	4	0.739 (25.34)
RFUG-067	101	110	9	0.996 (34.15)
Weighted Average			34	0.682 (23.38)
SOUTH RIB
RFUG-081	77	80	3	0.106 (3.63)
RFUG-082	80	83	3	0.065 (2.23)
RFUG-083	83	93	10	1.082 (37.10)
RFUG-084	93	96	3	0.894 (30.65)
RFUG-086	96	99	3	0.355 (12.17)
RFUG-087	99	107	8	0.028 (0.96)
RFUG-088	107	112	5	0.228 (7.82)
Weighted Average			35	0.470 (16.11)

Drilled By	# of Holes	Total Footage	Average Depth
Atna Resources-surface	88	68,069.6	774
Atna Resources-underground	48	15,349.0	320
HOMESTAKE	206	236,255.0	1,147
OTHER (PMC or Others)	2,119	1,072,320.0	395

Hole Number	Total Depth	RC Footage	Diamond Footage
APCX-201	440	0	440
APRF–202	788.5	0	788.5
APCX–203	845	400	445
APCX–204	980	385	595
APRF–205	674	385	289
APCX–206	222	0	222
APRF–207	635	635	0
APRF–208	1,090	700	390
APRF–209	676	350	326
APRF–210	645	400	245
APCX–211	1,010	500	510
APRF–212	980	600	380
APCX–213	1,100	700	400
APCX–214	1,020	600	420
APRF–215	1,010	600	410
APCX–216	1,045	460	585
APRF–217	920.5	600	320.5
APCX–218	1080	700	380
APCX–219	1,160	800	360
APCX–220	1248	970	278
APRF-221	1832	800	1032
APRF-222	650	650	0
APRF-223	1300	1000	300
APCX-224	1439	1060	379
APRF-225	1387	1080	307
APCX-226	1412	1100	312
APRF-227	780	260	520
APRF-228	500	500	0
APRF-229	432	432	0
APRF–229A	1515	1045	470
APRF–230	924.5	700	224.50
Total Drilled	29,740.5	18,412	11,328.5

Hole Number	Total Depth
UGRF-001	159
UGRF–002	373
UGRF-003	424
UGRF-004	409
UGRF-005	543
UGRF-006	638
UGRF-007	526
UGOG-001	191
UGOG-002	136
UGOG-003	102
UGOG-004	282
UGOG-005	422
UGOG-006	270
UGOG-007	183
UGOG-008	287
UGOG-009	213
UGOG-010	302
UGOG-011	418
UGOG-012	42
UGOG-013	263
UGOG-014	315
UGOG-015	213
UGOG-016	329
UGOG-017	539
UGOG-018	250
UGOG-019	133
UGOG-020	157
UGOG-021	162
UGOG-022	849
UGOG-023	187
UGOG-024	243
UGOG-025	450
UGOG-026	198
UGOG-027	203
UGOG-028	414
UGOG-029	170
UGOG-030	195
UGOG-031	198
UGOG-032	637
UGOG-033	667
UGOG-034	598
UGOG-035	452
UGCXW-001	163
UGCXW-002	204
UGCXW-003	178
UGCXW-004	207
UGCXW-005	357
UGCXW-006	498
Total Drilled	15,349

Character	Sample ID	Accepted Analytical Value	95% CI	Std Dev.
Oxide	OXE 21	0.651 ppm Au	+/- 0.012	0.026
	OXN 33	7.378 ppm Au	+/- 0.088	0.208
	OXK 18	3.463 ppm Au	+/- 0.058	0.132
Sulfide	SF 12	0.819 ppm Au	+/- 0.012	0.028
	SK 11	4.823 ppm Au	+/- 0.050	0.110
	SN 16	8.367 ppm Au, 17.64 ppm Ag	+/- 0.087	0.217
	SP 17	18.13 ppm Au	+/- 0.180	0.434
	SQ 18	30.49 ppm Au	+/- 0.350	0.88

Job	Sample	AU PPB	Au oz/ton	Accepted value	PPB Variance	Oz/ton for Standard	Oz/ton Variance
105-01-89	APCX-226 275	3609	0.12	3463	4.22%	0.10	16.83%
105-01-63	APRF-229 055	3438	0.12	3463	-0.72%	0.10	16.83%
105-00-49	APCX-224 335	3718	0.12	3463	7.36%	0.10	16.83%
105-00-31	APCX-216 215	6680	0.19	7378	-9.46%	0.22	-13.10%
104-29-42	APRF-227 035	20356	0.26	30490	-33.24%	0.89	-70.76%
104-28-94	APCX-224 215	18280	0.49	7378	147.76%	0.22	129.56%
104-27-01	APR-210 115	8190	0.29	7378	11.01%	0.22	33.83%
104-27-00	APCX-220 215	3900	0.11	3463	12.62%	0.10	10.89%
104-26-09	APRF-217 035	25821	0.77	30490	-15.31%	0.89	-13.64%
104-25-50	APCX-216 035	28550	0.69	30490	-6.36%	0.89	-22.86%
104-25-49	APRF-215 015	7000	0.21	8376	-16.43%	0.24	-14.86%
104-25-49	APRF-215 035	3690	0.11	3463	6.56%	0.10	10.89%
104-25-48	APCX-218 055	7765	0.24	7378	5.25%	0.22	10.60%
104-25-02	APCX-214 075	6596	0.21	8376	-21.25%	0.24	-14.86%
104-24-99	APR-211 115	3810	0.11	3463	10.02%	0.10	10.89%
104-24-60	APR-210 075	3541	0.10	7378	-52.01%	0.22	-52.60%
104-23-93	APR-208 155	7620	0.10	8376	-9.03%	0.24	-59.89%
104-23-91	APC-202 015	6720	0.19	7378	-8.92%	0.22	-10.78%

Decorative stone Initial analysis - September 9, 2004 BSI Inspectorate Final Report - Job No: 104-20-06 All Analyses by fire assay with AA finish
	Run 1		Run 2
Sample Number	Au ppb	Ag ppm	Au ppb		Combined Average	Clipped hi-lo
BWA-001	63	0.1	53		58	3
BWA-002	5	0.1	10		8	3
BWA-003	6	0.1	7		7	2
BWA-004	6	0.1	7		7	5
BWA-005	7	0.1	12		10	5
BWA-006	-5	-0.1	-5		3	3
BWA-007	7	0.1	6		7	7
BWA-008	-5	-0.1	-5		3	3
BWA-009	-5	-0.1	-5		3	3
BWA-010	6	-0.1	-5	2.5	4	4
BWB-001	6	-0.1	-5	2.5	4	4
BWB-002	8	0.1	-5	2.5	5	5
BWB-003	45	0.2	252		149	17
BWB-005	51	0.1	12		32	21
BWB-006	19	0.1	14		17	32
BWB-007	18	-0.1	24		21	58
BWB-009	16	0.1	-5	2.5	6
BWB-010	13	-0.1	10		12

Average	15		21		19	11
Standard Deviation	19		60		35	15
95% CI	9		27		16	7
Note: Less than detection values converted to +2.5 for statistical purposes

Grade Level	Number of samples	# Within 10% Error	# Within 20% Error	Rate below 10%	Rate Below 20%	Comparison Method
150 - 3000	110	74	21	67%	19%	AA:Grav
> 3000	140	120	136	86%	97%	Grav:Grav

Job Number	# Samples	# Failures	Rate
RE04074554	31	4	13%
RE04083801	62	4	6%
RE05000525	47	5	11%

Standard	Acceptable Gold Value (ppm)	95 pct Confidence limit	Standard Deviation
SQ18	30.49	0.35	0.88
SN16	8.376	0.087	0.217
SF12	0.819	0.012	0.028
OXN33	7.378	0.088	0.208
SP17	18.13	0.18	0.434
OXK18	3.463	0.058	0.132
OXE21	0.651	0.012	0.026
SK 11	4.823	0.05	0.11
OXL 25	5.852	0.048	0.105
OXH 29	1.298	0.015	0.033

Standard	Accepted Value (Au_ppb)	Lab_Certificate	Sample	AU_PPB	Comment
OXA4	81.1	06-330-00483-01	UGCXW-005 015	30	Blank insertion
SN16	8376	06-330-00046-01	APRF-245 075	10	Blank insertion
SN16	8376	05-330-00756-01	APRF-252 175	-5	Blank insertion
SF12	819	05-330-01193-02	APRF-258 135	31575	swapped std.
SF12	819	05-330-01400-01	UGOG-013 015	4589	Incorrect Std.
OXL 25	5852	06-330-00061-01	UGOG-018 075	30822	Incorrect Std.
OXK18	3463	05-330-01290-01	APRF-260 115	9452	Incorrect Std
SQ18	30490	06-330-00061-01	UGOG-018 035	6572	Incorrect Std.
SK 11	4823	05-330-01403-01	APRF-270 035	776	Incorrect Std.
SQ18	30490	05-330-01193-02	APRF-258 115	929	Swapped Std
SQ18	30490	05-330-00857-01	UGRF-002 095	870	Incorrect Std.
OXH 29	1298	05-330-00889-01	UGRF-003 015	30	Blank insertion
OXN33	7378	06-330-00266-02	UGRF-004 035	4178	Lab error
OXP32	14990	06-330-00507-01	UGCXW-006 115	7770	Incorrect Std.
OXP32	14990	06-330-00266-02	UGRF-004 075	10068	Lab error
SP17	18130	05-330-00581-01	APRF-241 115	13500	Lab error

Lab Certificate	Smp_ID	AU PPB	Accepted Value	Error Limit	Within Limits	Comment
105-08-90	APRF-231 005	50	7.8	36.6	No	Lab error
105-08-90	APRF-231 205	37	7.8	36.6	No	Marginal Failure
105-09-70	APRF-237 105	300	7.8	36.6	No	Lab error
05-330-01173-01	APRF-239 185	61	7.8	36.6	No	Lab error
05-330-01100-01	APRF-244 085	81	7.8	36.6	No	Lab error
105-12-67	APRF-245 005	130	7.8	36.6	No	Lab error
105-13-90	APRF-248 065	53	7.8	36.6	No	Lab error
105-13-90	APRF-248 105	57	7.8	36.6	No	Lab error
105-13-92	APRF-250 025	901	7.8	36.6	No	Lab error
105-13-92	APRF-250 045	232	7.8	36.6	No	Lab error
05-330-01099-01	APRF-251 145	77	7.8	36.6	No	Lab error
05-330-01099-01	APRF-251 165	38	7.8	36.6	No	Marginal Failure
05-330-01403-01	APRF-270 005	55	7.8	36.6	No	Lab error
05-330-01191-01	UGOG-003 025	73	7.8	36.6	No	Lab error
05-330-01172-01	UGOG-004 045	1005	7.8	36.6	No	Crusher Contam
05-330-01172-01	UGOG-004 065	663	7.8	36.6	No	Crusher Contam
05-330-01398-01	UGOG-006 025	83	7.8	36.6	No	Lab error
05-330-01293-01	UGOG-010 045	38	7.8	36.6	No	Marginal Failure
05-330-01293-01	UGOG-010 065	42	7.8	36.6	No	Marginal Failure
05-330-01399-01	UGOG-011 085	116	7.8	36.6	No	Crusher Contam
05-330-01400-01	UGOG-013 045	155	7.8	36.6	No	Crusher Contam
05-330-00889-03	UGRF-003 085	45	7.8	36.6	No	Marginal Failure

Lab_Certificate	HOLEID	FROM	TO	Duplicate Assay	Original Assay
06-330-00215-01	APRF-283	695	700	10	300
105-08-90	APRF-231	695	700	21	207
105-08-91	APRF-232	295	300	129	392
105-09-08	APRF-236	295	300	352	185
105-10-03	APRF-238	295	300	918	632
105-13-91	APRF-249	395	400	1089	1519
06-330-00067-01	APRF-275	395	400	1437	713
06-330-00262-01	APRF-285	495	500	1443	823
05-330-01340-01	APRF-267	695	700	3267	2056
06-330-00164-01	APRF-277	95	100	4630	3180

BSI Job #	Chemex Job # Check #1	HOLEID	FROM-TO	BSI (Au ppb)	Chemex Check #2 (Au ppb)	Chemex Check #1 Bad results (Au ppb)
105-08-92	RE05064671	APRF-233	520-525	31350	28900	22900
105-09-08	RE05064671	APRF-236	380-385	3734	3170	2700
105-13-90	RE05064671	APRF-248	360-365	4130	Insufficient pulp	493
105-13-90	RE05064671	APRF-248	390-395	12250	Insufficient pulp	7170
105-13-90	RE05064671	APRF-248	395-400	4030	Insufficient pulp	399
105-13-90	RE05064671	APRF-248	400-405	648	Insufficient pulp	4480
05-330-00621-01	RE05111829	APRF-232	708-710.9	17	<50	1890
05-330-00621-01	RE05111829	APRF-232	713.5-714.9	15	<50	3930
05-330-00378-01	RE05111829	APRF-237	700.2-706.4	4286	4220	340
05-330-00378-01	RE05111829	APRF-237	725-726	1798	1700	4030
05-330-00629-01	RE05111829	APRF-238	713-716.5	14	<50	3620
05-330-00629-01	RE05111829	APRF-238	724-727.5	2968	3040	1770
05-330-00629-01	RE05111829	APRF-238	727.5-729.3	371	420	1770
05-330-01173-01	RE05111829	APRF-239	613.7-619.2	22	<50	3370
05-330-00487-01	RE05111829	APRF-246	618.2-622	<5	<50	2380
105-13-89	RE05111829	APRF-247	570-575	12	<50	2230
05-330-01127-01	RE05111829	UGOG-001	32-35	4010	3690	5810
05-330-01127-01	RE05111829	UGOG-001	35-37	38895	42600	24800
05-330-01191-01	RE05111829	UGOG-003	53.3-57	53	<50	27900
05-330-01191-01	RE05111829	UGOG-003	57-59.4	22	<50	4410
05-330-01191-01	RE05111829	UGOG-003	62.9-67	102	<50	7680
05-330-01172-01	RE05111829	UGOG-004	109.5-114.5	7808	420	350
05-330-01172-01	RE05111829	UGOG-004	137-140.5	94700	140500	151500
05-330-01189-01	RE05111829	UGOG-005	27-28.8	129	100	3220
05-330-00857-01	RE05111829	UGRF-002	48-53	1511	1980	100

FIELD	Data Type
HOLEID	Hole name identifier
FROM	Beginning assay interval
TO	End assay interval
AUPPB	HMC-Chemex Fire assays in ppb
AUFA	HMC and Pinson Fire Assay in oz/ton
AUCN	HMC cyanide assays and Pinson AA cyanide soluble assays
AUMOZ	FACalc oz/ton multiplied by 1000
FACALC	Ounce per ton numbers used for modeling, Calculated from AUPPB or taken from AUFA fields
CN_FA	Ratio of CN vs. FA analyses
CN_FA_T	Unknown
CALC_CN	Unknown

Standard Name	Accepted Value and Error Range
CDN-GS-1	5.07 ± 0.43 g/t Au
CDN-GS-2	1.53 ± 0.18 g/t Au
CDN-GS-3	0.79 ± 0.07 g/t Au
CDN-GS-4	3.45 ± 0.21 g/t Au
CDN-GS-5	20.77 ± 0.91 g/t Au
CDN-GS-6	9.99 ± 0.50 g/t Au
CDN-GS-7	5.15 ± 0.46 g/t Au
CDN-GS-8	33.5 ± 1.7 g/t Au
CDN-GS-9	1.75 ± 0.14 g/t Au
CDN-GS-P3	0.30 ± 0.04 g/t Au
CDN-GS-10	0.82 ± 0.09 g/t Au
CDN-GS-1A	0.78 ± 0.08 g/t Au
CDN-GS-13	1.80 ± 0.18 g/t Au
CDN-GS-11	3.40 ± 0.27 g/t Au
CDN-GS-5A	5.10 ± 0.27 g/t Au
CDN-GS-14	7.47 ± 0.31 g/t Au
CDN-GS-12	9.98 ± 0.37 g/t Au
CDN-GS-15	15.31 ± 0.58 g/t Au
CDN-GS-20	20.60 ± 0.67 g/t Au
CDN-GS-30	33.5 ± 1.4 g/t Au

Samp_ID	New_Au_ppb	Accpt_Au_ppb	HighFlux_Au_PPM (Method1)	Grav_Au_PPM (Method 2)	Error tolerance +/-	Within Tolerance
RE04058845-27	770	780	0.75		+/- 0.140 g/t	No
RE04058845-5	4550	5150			+/- 0.460 g/t	Yes
RE04054309-14	740	820			+/- 0.09 g/t	Yes
RE04054309-74	9460	9990	9.65	9.59	+/- 0.50 g/t	Yes
RE04057300-344	4840	5150	5.12	4.73	+/- 0.460 g/t	Yes
RE04056499-24	16950	20770		20	+/- 0.91 g/t	Yes (Gravimetric value)
RE04056499-44	4770	5150	5.28	5.22	+/- 0.460 g/t	Yes
RE04056499-104	17950	20770	19.8	19.95	+/- 0.91 g/t	Yes

Stage 1 Metallurgical testing
Composite	Interval ft	Individual Samples
Composite	Interval ft	Individual Samples	Range Front Zone
RF_Met-1 (33941)	240-255 325-370	APRF-227: 054,056,057 APRF-228: 073,076,081,082
RF_Met-2 (33942)	683-719 668-684	APRF-202: 170,171,177,178 APRF-212: 159,162
RF_Met-4 (34259)	835-897 1172-1187 1466-1469.5 730-782	APRF-215: 200,201,202,217,218,219 APRF-225: 280,281,282,283 APRF-229A: 368,369 APRF-230: 172,173,174,190,192
OG-Met-1	35’ composite	35 ft (Left Rib) +32.5 ft (Right Rib) channel
Stage 2 Metallurgical testing
CX Zone
APCX-204	180-184	APC-204: 180-184
APCX-211	171-173	APC-211: 171-173
APCX-219	237-243	APCX-219: 237-243
APCX-226	266-270	APCX-226: 266-270
AMW-002	205-265	AMW-002: 047-059
Ogee Zone
OG004-1	109.5-147	UGOG-004: 034-042
OG004-2	150-192	UGOG-004: 043-054
OG004-3	192-230	UGOG-004: 055-067
OG004-4	230-257	UGOG-004: 068-076
OG010-2	156.5-184.8	UGOG-004: 048-058
OG010-3	184.8-215	UGOG-004: 059-070
OG010-1	50-65.9	UGOG-004: 012-016
OG013-1	86.8-118	UGOG-004: 034-046
OG013-2	213-227	UGOG-004: 074-078
OG015-2	74-101.7	UGOG-004: 023-036
OG015-3	101.7-135.5	UGOG-004: 037-052
OG017-1	319.5-348	UGOG-004: 087-102
OG017-2	349-375.5	UGOG-004: 103-114
OG017-3	375.5-401	UGOG-004: 115-122
OG017-4	459-492	UGOG-004: 141-152
OG018-1	12.8-25.2	UGOG-004: 006-008
OG018-2	40.4-62.7	UGOG-004: 016-022
OG019-1	78-105	UGOG-004: 022-032
OG021-1	76-110.8	UGOG-004: 022-037
OG022-1	659-684	UGOG-004: 157-168
OG022-2	733-748	UGOG-004: 180-188

Composite	ALI Fire g/t Au	CG			HLI			ALI
Composite	ALI Fire g/t Au	NaCN Soluble Au	% As	ppm Cu	% C(tot)	% C(C02)	% C(Org) *	% S(tot)	% S(804=)	% S= *
RF Composite Assays
RF_Met-1 (33941)	8.23	4.39	0.61	44	2.46	2.27	0.19	1.24	0.03	1.21
RF_Met-2 (33942)	14.74	8.98	1.01	36	1.93	1.76	0.17	3.41	0.80	2.61
RF_Met-4 (34259)	14.74	13.1	2.02	38	2.69	2.43	0.26	2.40	0.08	2.32
CX Composite Assays
APCX-204	8.16	7.7	0.066	36	5.46	5.34	0.12	<0.03	<0.03	-
APCX-211	8.33	7.1	0.13	19	4.47	4.29	0.18	<0.03	<0.03	-
APCX-219	10.25	6.1	0.18	30	1.04	0.77	0.27	0.92	0.08	0.84
APCX-226	17.50	7.4	1.65	74	3.45	2.7	0.75	1.53	<0.03	1.53
AMW-002 (P2895 B)	10.29	7.9	0.054	179	3.43	3.08	0.35	0.07	0.04	0.03
Ogee Composite Assays
UGOG-004-1	26.4	26.9	0.27	7.6	0.69	0.39	0.30	0.86	0.08	0.78
UGOG-004-2	30.9	19.2	0.27	8.5	0.29	<0.02	0.29	1.16	0.08	1.08
UGOG-004-3	41.8	31.6	0.20	8.4	0.18	<0.02	0.18	1.47	0.13	1.34
UGOG-004-4	22.3	20.3	0.09	5.8	0.53	0.35	0.18	0.42	<0.03	0.42
UGOG-010-1	45.3	24.8	0.36	6.7	0.54	<0.02	0.54	1.87	<0.03	1.87
UGOG-010-2	54.7	49.8	0.24	6.3	0.32	<0.02	0.32	0.11	<0.03	0.11
UGOG-010-3	41.6	35.9	0.14	7.1	0.21	<0.02	0.21	0.05	<0.03	0.05
UGOG-013-1	13.2	5.59	0.20	4.9	1.20	0.86	0.34	1.64	<0.03	1.64
UGOG-013-2	23.8	22.2	0.13	5.7	2.68	2.46	0.22	0.03	<0.03	0.03
UGOG-015-2	9.94	6.05	0.15	4.6	0.92	0.68	0.24	1.00	<0.03	1.00
UGOG-015-3	27.6	12.1	0.36	9.4	0.32	<0.02	0.32	2.44	<0.03	2.44
UGOG-017-1	28.4	3.08	0.46	6.3	0.90	0.59	0.31	2.75	0.04	2.71
UGOG-017-2	23.0	16.5	0.14	8.8	0.58	0.19	0.39	1.37	<0.03	1.37
UGOG-017-3	45.3	29.9	0.14	10.8	0.57	0.40	0.17	0.55	<0.03	0.55
UGOG-017-4	16.1	11.1	0.12	6.1	0.74	0.51	0.23	0.95	<0.03	0.95
UGOG-018-1	15.8	12.3	0.08	5.1	1.40	0.96	0.44	<0.03	<0.03	<0.03
UGOG-018-2	14.2	11.7	0.11	7.9	0.74	0.09	0.65	0.04	<0.03	0.04
UGOG-019-1	43.8	34.1	0.24	7.1	0.28	0.02	0.26	0.45	<0.03	0.45
UGOG-021-1	24.8	20.2	0.21	5.5	0.36	<0.02	0.36	0.08	0.03	0.05
UGOG-022-1	23.5	4.21	0.34	2.9	2.18	2.12	0.06	1.48	<0.03	1.48
UGOG-022-2	7.06	5.19	0.14	1.7	3.69	3.65	<0.05	0.08	0.03	0.05
OG-Met-1 (LR&RR)	13.71	11.4	0.47	44	1.11	0.82	0.29	0.04	0.04	<0.02

Composite	Head Grade	Shake Leach Extraction Grade (SLE)	% SLE	Autoclave Grade g/mt	Autoclave %
RF Zone
RF_Met-1 (33941)	8.23	4.39	53	8.38	93.0
RF_Met-2 (33942)	14.74	8.98	61	14.88	95.2
RF_Met-4 (34259)	14.74	1.61	11	13.1	88.5
CX Zone
APCX-219	10.25	6.1	60	7.6	91.0
APCX-226	17.50	7.4	42	14.84	93.8
Ogee Zone
OG-Met-1 (LR&RR)	13.72	11.79	86	11.4	93.2

Impurities	Range
Sulfur: total sulfide	1.2 to 3.4%
Carbon: total organic carbon	0.17 to 0.29%
Carbonate Level	6.4 to 12.7%
As (arsenic)	1000 to 6000 ppm
Sb (antimony)	85 to 300 ppm
Pb (lead)	7 to 16 ppm
Zn (zinc)	58 to 160 ppm
Cu (copper)	29 to 58 ppm
Mn (manganese)	40-275 ppm
Hg (mercury)	60 to 100+ ppb

Composites	Material type		NaCN/ FA %	HNO₃/ NaCN/ FA %	TOC
Composites	Visual	Test	NaCN/ FA %	HNO₃/ NaCN/ FA %	TOC

Range Front Zone
RF_Met-1 (33941)	Oxide	Autoclave - Leach	53	95	0.19
RF_Met-2 (33942)	Sulfide	Autoclave - Leach	61	99	0.17
RF_Met-4 (34259)	Sulfide	Direct Leach	89	88	0.26
CX Zone
APCX-204	Sulfide	Direct Leach	94	95	0.12
APCX-211	Sulfide	Direct Leach	85	96	0.18
APCX-219	Sulfide	Autoclave - Leach	60	89	0.27
APCX-226	Sulfide	Autoclave - Leach	42	81	0.75
AMW-002 (P2895 B)	Oxide	Autoclave - Leach	77	93	0.35
Ogee Zone
OG004-1	Mix	Direct Leach	102		0.3
OG004-2	Mix	Autoclave - Leach	62		0.29
OG004-3	Mix	Autoclave - Leach	76		0.18
OG004-4	Mix	Direct Leach	91		0.18
OG010-2	Mix	Autoclave - Leach	55		0.54
OG010-3	Mix	Direct Leach	91		0.32
OG010-1	Oxide	Direct Leach	86		0.21
OG013-1	Oxide	Autoclave - Leach	42		0.34
OG013-2	Mix	Direct Leach	93		0.22
OG015-2	Mix	Autoclave - Leach	61		0.24
OG015-3	Mix	Autoclave - Leach	44		0.32
OG017-1	Mix	Autoclave - Leach	11		0.31
OG017-2	Mix	Autoclave - Leach	72		0.39
OG017-3	Mix	Autoclave - Leach	66		0.17
OG017-4	Mix	Autoclave - Leach	69		0.23
OG018-1	Oxide	Autoclave - Leach	78		0.44
OG018-2	Oxide	Autoclave - Leach	82		0.65
OG019-1	Oxide	Autoclave - Leach	78		0.26
OG021-1	Oxide	Autoclave - Leach	81		0.36
OG022-1	Sulfide	Autoclave - Leach	18		0.06
OG022-2	Sulfide	Autoclave - Leach	74		<0.05
OG-Met-1 (RR+LR)	Oxide	Autoclave - Leach	83	93	0.29

*Domain*	*Code#*	*Description*
CX Fault	ZONE=1	CX Fault zone
RF Fault	ZONE=2	Range Front Fault zone
CX HG Zone	HGZNE=1	Portion of CX fault>0.1 opt Au
CX LG Zone	HGZNE=2	Portion of CX fault <0.1 opt Au
RF HG Zone	HGZNE=3	Portion of Range Front fault>0.1 opt Au
RF HG Zone	HGZNE=4	Portion of Range Front fault <0.1 opt Au

*Gold (opt)*	*CX Fault*	*Range Front Fault*
Total #/ft samples	6,750/33,661	2,665/13,269
Min	0.0001	0.0001
Max	1.840	2.337
Mean	0.030	0.037
Std Dev	0.095	0.122