TEXT

TABLE OF CONTENTS

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1.0	Summary	1
1.1	Introduction	1
1.2	Property Exploration	2
1.3	Geological Setting	2
1.4	Xxxxxx Project - Geology, Alteration, and Mineralization	3
1.5	Mineral Resource Estimation	3
1.6	Conclusions	5
1.7	Recommendations	6
2.0	Introduction and Terms of Reference	8
2.1	Introduction	8
2.2	Terms of Reference	8
3.0	Disclaimer	9
4.0	Property Description and Location	10
4.1	Location	10
4.2	Land Status and Agreements	11
4.2.1	Xxxxxx Mining Company Agreement	12
4.2.2	Underlying Agreements - Unpatented Federal Mining Claims	15
4.2.2.1	BEE DEE group	15
4.2.2.2	X. X. Xxxxx Unpatented Claims	15
4.2.2.3	VEK/Xxxxxx Unpatented Claims	15
4.2.3	Underlying Agreements - Patented Fee Lands	16
4.2.3.1	Xxxxxxxxxx, et al Lease	16
4.2.3.2	X. Xxxxx Lease	16
4.2.3.3	Xxxxxx Lease	16
4.2.4	Underlying Agreements - Other Agreements	16
4.2.4.1	Xxxxxxxxxx - Seven Dot Agreement	17
4.2.4.2	Xxxxxxx (Glamis Gold) Acquisition	17
4.2.4.3	Cordilleran Overriding Royalty	18
4.2.4.4	Xxxxxxxx Gold/Newmont Capital Overriding Royalty	18
4.2.4.5	Echo Bay Lands	18
4.3	Royalty Summary	18
5.0	Accessibility, Climate, Local Resources, Infrastructure and Physiography	20
5.1	Accessibility	20
5.2	Climate	21
5.3	Local Resources and Infrastructure	21
5.3.1	Background	21
5.3.2	Physical Infrastructure	22
5.4	Adjacent Operations	22
5.5	Physiography, Flora, and Fauna	22
5.5.1	Physiography	22
5.5.2	Flora	22
5.5.3	Fauna	22
6.0	History	24
6.1	Summary	24
6.2	Initial Discovery	24
6.3	Pre-1970 Development	24
6.4	Cordex Syndicate Exploration and Development	24
6.5	Xxxxxx Mining Company Exploration and Development	25
6.6	Homestake-Barrick Exploration	25
6.7	Xxxxxx Project Production Summary	25
6.8	Atna Resources Exploration	25
7.0	Geological Setting	26
7.1	Regional Setting	26
7.2	Xxxxxx Mine Setting	30
8.0	Deposit Type	34
8.1	Sediment-hosted, Xxxxxx-type Gold System	34
9.0	Mineralization	36
9.1	Host Rocks	36
9.2	Alteration	37
9.3	Mineralized Zone Configuration	38
10.0	Exploration	42
10.1	Introduction	42
10.2	Geologic Mapping and Geochemical Sampling	42
10.3	Drilling	43
10.4	Trenching and Channel Sampling	43
10.5	Geophysics	43
10.6	Underground Drifting / Evaluation	44
11.0	Drilling	45
11.1	Summary of Past and Present Programs	45
11.1.1	Drilling by Earlier Operators	45
11.1.2	Drilling by Atna Resources	45
11.2	Drilling methods	46
11.2.1	Reverse Circulation Rotary Drilling	47
11.2.2	Diamond Core Drilling	48
11.3	Logging	48
11.3.1	Reverse Circulation Rotary Chip Logging	48
11.3.2	Diamond Drill Core Logging	48
12.0	Sampling Method and Approach	50
12.1	Sampling Methods	50
12.1.1	Reverse Circulation Rotary	50
12.1.2	Diamond Core Drilling	50
12.2	Sample Quality - Recovery	50
12.2.1	Reverse Circulation Rotary	50
12.2.2	Core	50
12.3	Sample Interval	51
12.3.1	Reverse Circulation Rotary	51
12.3.2	Core	51
12.4	Sample Preparation, Quality Control Measures and Security	51
12.4.1	Sample Preparation and Quality Control Measures - RC Rotary Drilling	51
12.4.2	Sample Preparation and Quality Control Measures - Core Drilling	52
12.4.3	Security - Reverse Circulation and Core Samples	54
12.4.4.1	On-site	54
12.4.4.2	Laboratory Sample Preparation	54
12.5	Certified Standard Insertion	55
12.5.1	Protocol	55
12.5.2	Summary of Results	56
12.6	Blank Sample Insertion	59
12.6.1	Protocol	59
12.6.2	Summary of Results	59
12.7	Duplicate Samples - Reverse Circulation Rotary	60
12.7.1	Protocol	60
12.7.2	Summary of Results	61
12.8	Check Assays of Mineralized Samples	62
12.8.1	Protocol	62
12.8.2	Summary of Results	62
12.9	Laboratory Quality Assurance and Control	66
13.0	Data Verification	68
13.1	Summary	68
13.2	Database of Previous Drilling	68
13.2.1	Tables	69
13.2.2	Data Corrections	69
13.2.3	General description of pre Atna assay results and procedures.	71
13.3	Pulp Selection	72
13.4	Re-Assay Methods	73
13.4.1	Processing by BSi Inspectorate	73
13.4.2	Processing by ALS Chemex	73
13.5	Re-Assay Results	74
13.6	Discussion	76
14.0	Database Audit	78
14.1	Analytical Data	78
14.1.1	Validation Process	78
14.2	Drill-hole Collar Location	79
15.0	Adjacent Properties	80
15.1	Summary	80
15.2	Xxxxxx Mine	80
15.3.	Xxxxxxxx/Turquoise Ridge Mine Complex	80
15.3.1	History	80
15.3.2	Geologic Setting	81
15.3.3	Gold geology	81
15.4	Twin Creeks	82
15.4.1	Physiography	82
15.4.2	Stratigraphy	83
15.4.3	Structure	83
15.4.4	Erosion & Alluviation	84
15.4.5	Mineralization	84
15.4.6	Alteration	85
15.4.7	Summary	86
16.0	Mineral Processing and Metallurgical Testing	87
17.0	Mineral Resources Estimates	88
17.1	Introduction	88
17.2	Geologic Model, Domains and Coding	89
17.2.1	Geologic Model	89
17.2.2	Domains and Coding	89
17.3	Available Data	94
17.4	Compositing	96
17.5	Exploratory Data Analysis	96
17.5.1	Basic Statistics by Domain	97
17.5.2	Contact Profiles	98
17.5.3	Histograms and Log-Probability Plots	99
17.5.4	Conclusions and Modeling Implications	103
17.6	Bulk Density Data	103
17.7	Evaluation of Outlier Grades	103
17.8	Variography	104
17.9	Model Setup and Limits	105
17.10	Interpolation Parameters	107
17.11	Validation	108
17.11.1	Visual Inspection	108
17.11.2	Model Checks for Change of Support	109
17.11.3	Comparison of Interpolation Methods	111
17.11.4	Swath Plots	112
17.12	Resource Classification	114
17.13	Key Assumptions of Economic and Technical Parameters	117
17.14	Mineral Resources	118
18.0	Other Relevant Data and Information	121
19.0	Interpretation and Conclusions	123
20.0	Recommendations	124
20.1	Recommended Program and Budget	125
	References	128

FIGURES

Figure 4-1: Location Map of the Xxxxxx Mine Project	11
Figure 4-2: Generalized Land Status Map	12
Figure 4-3: Detailed Land Status and Agreement Map	14
Figure 7-1: Regional Stratigraphy	27
Figure 7-2: Regional Geology	29
Figure 7-3: TectonoStratigraphy	30
Figure 7-4: Mine Site Geology	33
Figure 9-1: Grade-Thickness Long Section, CX Fault Zone	39
Figure 9-2: Grade-Thickness Long Section, Range Front Fault Zone	41
Figure 12-1:  Standards Run With Gravimetric Finish.	58
Figure 12-2: Standards Run With AA Finish	59
Figure 12-3: Comparison of Duplicate Samples Taken at the Drill vs. Assay Sample.	62
Figure 12-4: Sample Variance of All Check Samples	63
Figure 12-5: X-Y Pair Plot of Check vs. Original Assay, By Grade	64
Figure 12-6: Variance of AA-Finish samples < 3000 ppb	65
Figure 12-7: Gravimetric Finish Comparisons - -Check Assays to Original Assays	66
Figure 12-8: Laboratory Blind Duplicates	67
Figure 13-1: Test Result Variances	75
Figure 17-1: Rotated Vertical NW-SE Cross-Sections	88
Figure 17-2: Cross Sectional View of CX and Range Front Fault Zones	90
Figure 17-3: Isometric View of CX and Range Front Fault Zones	91
Figure 17-4: Sectional View of HG Zone Trends Within Fault Zones	92
Figure 17-5: Comparison of Indicator Probability Estimation Techniques - Plan	93
Figure 17-6: Comparison of Indicator Probability Estimation Techniques - Section	93
Figure 17-7: CX High-Grade Zone Modifications	94
Figure 17-8: Range Front HG Zone Modifications	94
Figure 17-9: Distribution of Drill Holes by Type - CX Zone	95
Figure 17-10: Distribution of Drill Holes by Type - Range Front Zone	96
Figure 17-11: Contact Profile of CX HG vs. LG Domains	99
Figure 17-12: Contact Profile of Range Front HG vs. LG Domains	99
Figure 17-13: Histogram of Gold in CX Fault Zone	100
Figure 17-14: Histogram of Gold in CX HG Zone	100
Figure 17-15: Histogram of Gold in CX LG Portion of Fault Zone	101
Figure 17-16: Histogram of Gold in Range Front Fault Zone	101
Figure 17-17: Histogram of Gold in Range Front HG Zone	102
Figure 17-18: Histogram of Gold in Range Front LG Portion of Fault Zone	102
Figure 17-19: Log-Probability Plot of Gold in CX Fault Zone	102
Figure 17-20: Log-Probability Plot of Gold in CX HG Zone	103
Figure 17-21: Log-Probability Plot of Gold in CX LG Portion of Fault Zone	103
Figure 17-22: Log-Probability Plot of Gold in Range Front Fault Zone	104
Figure 17-23: Log-Probability Plot of Gold in Range Front HG Zone	104
Figure 17-24: Log-Probability Plot of Gold in Range Front LG Portion of Fault Zone	105
Figure 17-25: CX Zone Block Model Limits	108
Figure 17-26: Range Front Zone Block Model Limits	109
Figure 17-27: CX HG Zone Drill Hole and Block Grade Distribution	112
Figure 17-28: Range Front HG Zone Drill Hole and Block Grade Distribution	112
Figure 17-29: CX HG Zone Krige/IDW/Herco Plots	112
Figure 17-30: Range Front HG Zone Krige/IDW/Herco Plots	113
Figure 17-31: Swath Plot CX HG Zone, East-West	114
Figure 17-32: Swath Plots CX HG Zone, North-South	114
Figure 17-33: Swath Plot CX HG Zone, Vertical	115
Figure 17-34: Swath Plot Range Front HG Zone, East-West	115
Figure 17-35: Swath Plot Range Front HG Zone, North-South	115
Figure 17-36: Swath Plot Range Front HG Zone, Vertical	116
Figure 17-37: Confidence Limits Distribution by Drill-hole Spacing	117
Figure 17-38: CX Zone Resource Classification	119
Figure 17-39: Range Front Zone Resource Classification	119
Figure 18-1: Location of Final Phase I Range Front Holes (APRF-229A and APRF-230)	123

TABLES

Table 1-1: Mineral Resource, Combined CX and Range Front Zones	5
Table 4-1: Royalty Summary	19
Table 5-1: Seasonal Temperature Variations	21
Table 9-1: Inventory of Alteration Types in Atna Phase I Drilling	38
Table 11-1: Summary of Previous Drilling.	45
Table 11-2: Summary of Atna Resources Phase I Drilling	46
Table 12-1: RockLabs Reference Material.	55
Table 12-2: Failed Gravimetric standards	57
Table 12-3: Decorative Stone (Blank Sample) Analysis	60
Table 12-4: Statistics of Lab Variance in AA vs. Gravimetric Finish	64
Table 12-5: Failure Rate for 3 Check Assay Jobs.	66
Table 13-1: Field Definitions in (HMC) Xxxxxx Database	70
Table 13-2: CDN Resource Laboratory Standards	73
Table 13-3: Analyses of Standards Used in Testing.	75
Table 17-1: Summary of Geologic Domains	94
Table 17-2: Drill-hole Sample Statistics by Fault Zone	97
Table 17-3: Drill-hole Sample Statistics by HG/LG Portion of Fault Zone	98
Table 17-4: Proportion of Contained Gold in Decile/Percentile of Samples	106
Table 17-5: Variogram Parameters	107
Table 17-6: CX Zone Block Model Limits	108
Table 17-7: Range Front Zone Block Model Limits	109
Table 17-8: Interpolation Parameters for Ordinary Krige Estimates	110
Table 17-9: Interpolation Parameters for IDW Estimates	110
Table 17-10: Comparison of Interpolation Methods	113
Table 17-11: Quarterly and Yearly Confidence Limits Determination	117
Table 17-12: Mineral Resource, CX Zone	120
Table 17-13: Mineral Resource, Range Front Zone	121
Table 17-14: Mineral Resource, Combined CX and Range Front Zones	122
Table 18-1: Significant Drill Results From Holes APRF-229A and APRF-230	123

APPENDICES

APPENDIX - SECTION 4: Property Description and Location

Appendix 4-1, Xxxxxx 2005 Land Budget Projection

Appendix 4-2, Exhibit A, Part I, Unpatented Mining Claims

Appendix 4-3, Exhibit A, Part II, Fee Lands

APPENDIX - SECTION 6: History

Appendix 6-1, Discovery & Production History

APPENDIX - SECTION 9: Mineralization

Appendix 9-1 Section 6800

Appendix 9-2 Section 7100

Appendix 9-3 Section 7700

APPENDIX - SECTION 12: Sampling Methods and Approach

Appendix 12-1, Sample Sequence

Appendix 12-2 OXE21

Appendix 12-3 OXN33

Appendix 12-4 OXK18

Appendix 12-5 SF12

Appendix 12-6 SK11

Appendix 12-7 SN16

Appendix 12-8 SP17

Appendix 12-9 SQ18

Appendix 12-10 CHECKS and STANDARDS

APPENDIX - SECTION 13: Data Verification

Appendix 13-1, Pulp Re-Assay List

APPENDIX - SECTION 14: Database Audit

Appendix 14-1 Error Analysis of Randomly Selected Xxxxxx Database Assays

Appendix 14-2, Drill Hole Collar Table

APPENDIX - SECTION 17: Mineral Resources Estimates

Appendix 17-1, CX HG Zone Correlogram (pdf file)

Appendix 17-2, Range Front HG Zone Correlogram (pdf file)

Appendix 17-3, Block Model and Drill Hole Cross Sections (pdf file)

1.0 Summary

1.1   Introduction

In the fall of 2004, Atna Resources commissioned Xxxxxx Xxx, P. Xxx. of Sim Geological, to prepare an independent Qualified Person’s mineral resource estimate, review and Technical Report for the Xxxxxx Mine Property located in Humboldt County, Nevada. Xxxxxx Xxx is an independent consultant specializing in mineral resource estimations. Mr. Xxx received assistance in the preparation of this report from several of Atna’s personnel including Xxxxxxx Xxxxxxx, L.G., Vice-President, Exploration and site project geologists, Xxxx Xxxxxxx, Xxxxxx XxxXxxxxx and Xxxxx (Xxx) Xxxxxxx. Xxxxxx Xxx, P.Xxx serves as the Qualified Person for the preparation of this Technical Report as defined in National Instrument 43-101, Standards of Disclosure for Mineral Project, and in compliance with Form 43-101F1 (the Technical Report). Mr. Xxx understands that this report will be submitted to the TSX Exchange in support of the required filings by Atna Resources Ltd.

This report was initially released under an effective date of February 15, 2005. It has been amended in December 2005 in order to correct several disclosure and filing issues. There have not been any changes to the overall content or conclusions of the report with respect to data collected subsequent to the original release of the report. Therefore, the effective date of this report remains February 15, 2005.

This report represents the first Technical Report completed on the Xxxxxx Mine Property for Atna Resources Ltd.

Work reported on in this Technical Report includes the review of a exploration programs at the property conducted prior to Atna Resources involvement, including, but not limited to: exploration drilling campaigns carried out by the Cordex I Syndicate, Xxxxxx Mining Company, Homestake Mining Company (as operator of the property), and Xxxxxxx Gold Exploration (subsequent to Xxxxxxx’x acquisition of Homestake Mining). Additionally, this Technical report summarizes all data collected by Atna Resources as part of its Phase 1 exploration drilling efforts between August 2004 and February 15^th, 2005.

The Xxxxxx Property is located approximately 27 air miles northeast of the town of Winnemucca in north-central Nevada. Atna controls, through its agreement with Xxxxxx Mining Company, approximately 12,785 acres of fee lands and unpatented federal lode claims centered on the Xxxxxx Mine (UTM coordinates: 478294.7E, 4553517.9N, Zone 11, North American Datum of 1927). Access to the property is superb, with paved interstate and secondary state highways to within four (4) miles of the property, and well maintained mine-access, gravel roads traversing the property.

The property is comprised of 568 federal unpatented mining claims (8,485 acres) and 4,280 acres of patented fee lands that are either owned outright or leased by Xxxxxx Mining Company and subject to the Atna Resources agreement.

Atna’s rights to the property are governed by an Exploration Agreement with Option to Venture with Xxxxxx Mining Company, a Nevada Partnership dated August 12, 2004. Xxxxxx Mining Company is now 100% owned by Xxxxxxx Gold as the result of Xxxxxxx’x acquisition of Homestake Mining Company, who was a 50% co-owner of Xxxxxx Mining Company. To earn a 70% interest in the property, Atna must complete the expenditure of US$12.0 million on qualifying exploration expenditures over a four (4) year period, including US$1.5 million in the first 12 months of the agreement (completed) and deliver to Xxxxxx Mining Company a pre-feasibility study recommending further development of the property. Xxxx has spent in excess of US$2.0 million on the property through February 15, 2005. After completion of Atna’s earn-in requirements, Xxxxxx Mining Company may elect to 1) back into the project to a 70% joint venture interest through the expenditure of US$30 million over a three (3) year period (thus leaving Atna with a 30% interest); 2) participate in the project’s continued development at the 70% Atna / 30% Xxxxxx Mining Company interest levels; or 3) sell Atna its 30% interest in the property for US$15.0 million.

The Xxxxxx Mine produced approximately 985,000 ounces of gold from several open pit mines from the 1970s through mine closure in 2000. Mineralization is hosted dominantly by black calcareous and carbonaceous silty shales, and limy siltstones of the lower portion of the Ordovician Comus Formation. Like the neighboring active mines of Xxxxxxxx, Turquoise Ridge, and the Twin Creeks complex, mineralization at the property may be characterized as being a "Xxxxxx-type," sediment-hosted gold system. Low-grade mineralization was exploited by open pit methods which have been mostly exhausted. However, gold mineralization continues beneath the pit limits along zones of strong structural preparation that served as “feeder-faults” for mineralizing fluids to the low grade zones historically mined at Xxxxxx.

1.2   Property Exploration

Exploration has been carried out at the property from the late 1930’s through to the present by a number of individuals and mining/exploration companies. The original discovery on the property was made by Xxxxxx Xxxxxx and Xxxxxxx Xxxx in the mid to late-1930s, but production from their gold discovery in sedimentary rocks did not produce until after World War II. Ore from this original discovery was shipped and processed at the Xxxxxxxx mine mill in 1949 and 1950 from a small open pit. Total production by Xxxxxx and Ogee amounted to approximately 100,000 tons grading approximately 0.14 opt gold.

Minor exploration occurred on the property between 1950 and 1970 with a few drill holes completed by Homestake Mining and Nevada Goldfields. The property was dormant until the Cordex Syndicate I (Xxxx Xxxxxxxxx, Xxxxx Xxxxx, and Rayrock Mines) began picking up property in the district in 1970 and began exploration drilling shortly thereafter. The Cordex and/or Rayrock work resulted in the discovery of the A, B, C, CX, CX West, and Mag deposits at the property between 1970 and 1985. Production from these deposits generated approximately 986,000 ounces of gold up until the project was closed in 2000. Each of the operators conducted extensive exploration including over 2,000 exploration and development drill holes, geophysical and geochemical surveys as well as geologic mapping and sampling. The last major exploration effort was conducted from 1997 through 2000 by Homestake (on behalf of the Homestake-Barrick partnership, PMC). Homestake drilled over 200 holes in numerous targets throughout the property with the hope of extending the mine life via a new discovery. This work confirmed that the gold mineralization within the CX fault zone continued below and beyond the existing pit limits and that mineralization associated with the Range Front fault zones had significant strike and down-dip continuity.

Atna Resources’ work has focused on the exploration and confirmation of the CX and Range Front mineral zones. Atna has completed 31 drill holes totaling 29,740.5 feet. Of the 31 drill holes completed, at the cut-off date for the mineral resource section of the report 29 holes had final laboratory assay results. Two holes remained pending (APRF-229A and APRF-230).

1.3   Geological Setting

Sediment-hosted, "Xxxxxx-type" gold systems account for the vast majority of the gold production in Nevada and for most economic gold discoveries made since 1975. Mineralization in these environments lies mainly in four geographic belts of mostly Paleozoic carbonate rocks. These belts are located in north-central Nevada, and the three most productive pass through the townsites of Carlin (“Xxxxxx Trend”), Battle Mountain (“Battle Mountain-Eureka Trend”) and Golconda (“Xxxxxxxx Trend”). The fourth belt, the “Independence Trend”, is located north of the town of Elko and is the location of the Jerritt Canyon group of mines (Queenstake Resources) and the Big Springs Mine (Golden Gate Resources). Collectively, these belts hold a geochemical endowment of over 200 million ounces of gold.

Northern Nevada’s geology is complex, but is dominated by Basin and Range brittle, extensional tectonics resulting in the characteristic northerly-trending basins and ranges throughout the state. In the ranges (xxxxx blocks), marine sedimentary rocks are exposed spanning the entire range of Paleozoic rocks that are complexly thrust, intruded by younger Cretaceous intrusions. and both intruded and capped by Tertiary-age volcanic rocks. Of primary interest to the Xxxxxx mineral system are the lower Ordovician rocks of the Comus Formation and the underlying Cambrian Xxxxxx Formation.

1.4   Xxxxxx Project - Geology, Alteration, and Mineralization

Xxxxxx’x geologic setting is dominated by three main elements - sedimentary stratigraphy of Ordovician to Cambrian age, Cretaceous plutonism, and high angle, brittle structural deformation. Marine sedimentary rocks make up the bulk of the rocks within the property sequence and serve as the main host rocks for the gold mineral system. The sediments were intruded in the Cretaceous by granodiorite intrusions including the Xxxxxx Mountain stock (which forms the footwall of the Range Front mineral zone) and created a broad contact metamorphic aureole into the sedimentary sequence at the Xxxxxx Mine project. Tertiary and older brittle fault structures along the southeastern flank of the Xxxxxx Mountain Range create a network of permeable fractures that provided primary access routes for gold mineralizing hydrothermal solutions into the Ordovician and Cambrian host rocks.

Gold mineralization at the project is characterized as a Xxxxxx-type, sediment-hosted system similar to most other productive gold systems of this type in Northern Nevada including the adjacent properties in commercial production. These gold systems are associated with local silicification of the sediments, particularly along structural feeder faults, broad zones of de-calcification of the sedimentary section, and the introduction of very fine-grained pyrite, xxxxxxxx-pyrite, orpiment, realgar, stibnite and, commonly, re-mobilized carbon. Mineralization is often found in both structurally-prepared zones along faults and within locally important receptive host lithologies (e.g. thin-bedded calcareous carbonaceous siltstones, debris flows, and karst breccias). Mineralization may be associated with argillization of the host lithologies and the development of sericite in the matrix of the clastic sedimentary rocks.

Gold is primarily associated with the introduction of fine-grained disseminated auriferous and xxxxxxxx pyrite. Additionally, minor silicification may also be present as a late-stage flooding of brecciated limy sedimentary rocks creating “jasperoid” bodies along the structural traces. Two principal zones of mineralization are the focus of this Technical Report - the CX zone and the Range Front zone. The CX zone was the principal fault zone controlling mineralization previously mined in the CX pit; the Range Front zone has not been mined historically and the initial zone of mineralization was outlined by Homestake’s work in the late 1990s. Both of these zones are controlled principally by high-angle brittle fault zones. However, another important mineral control is the presence of receptive host lithologies along the fault zones.

1.5   Mineral Resource Estimation

The mineral resource estimations were made from 3-dimensional block models generated using commercial mine planning software (MineSight).

The mineral resource estimate was generated from drill-hole sample assay results and a geologic model which relies on the spatial distribution of gold. Individual domains, reflecting distinct zones or types of mineralization, have been defined and interpolation characteristics have been defined for each domain based on the geology, drill-hole spacing and geostatistical analysis of the data. The degree of confidence in the resources has been classified based on the proximity to sample locations and/or surface exposures and are reported, as required by NI 43-101, according to the CIM standards on Mineral Resources and Reserves.

The report includes only estimates for mineral resource. No mineral reserves are prepared or reported in this Technical Report.

Based on the information available in the drilling database (401 drill holes including core and RC holes), 3-dimensional wireframe solids were generated for the CX and Range Front fault zones. The shape and location of the gold bearing “high-grade” (HG) portion within these fault zones were estimated using a combination of geostatistical and geologically interpretative methods.

The block model grade interpolation, by ordinary kriging, was conducted using hard boundary code matching within the HG zone domains. For comparison purposes, inverse distance and nearest neighbor estimations of grade were also conducted. The Mineral Resources contained within the HG zones of the combined CX and Range Front zones are shown in Table 1-1 and are classified as Measured, Indicated and Inferred as per CIM reporting standards.

Based on projected economic and technical parameters, derived primarily from nearby (similar) operations, the “base case” operating cut-off grade for the Xxxxxx deposit in estimated to be 0.20 ounces per ton gold.

Table 1-1: 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 +	0.15	1,758	0.30	519
Indicated	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 = 0.20optAu)

1.6   Conclusions

Mineralization located within the CX and Range Front mineralized zones at the Xxxxxx Mine project represent significant bodies of gold-mineralized rocks with characteristics similar to the economic and productive Xxxxxx-type gold systems currently in production in the Xxxxxxxx Gold Belt and other districts in Northern Nevada.

Gold mineralization is hosted by the same stratigraphic units hosting the Xxxxxxxx deposit. Xxxxxx’x mineralization has a similar structural control to that present at the Xxxxxxxx Mine, five miles to the north, occurring within the network of faults making up the xxxxx bounding, Basin and Range fault zone along the southeastern margin of the Xxxxxx Mountains. Gold grades within both the CX 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 31 drill holes, the geologic understanding of the mineral system’s configuration, structure and stratigraphic control are adequate to support the current resource model contained within this Technical Report.

Generally, drilling is rather wide spread, particularly in the deeper portions of the two gold mineralized zones that are the focus of Atna’s work. Additional drilling planned in conjunction with underground access development will provide the additional data density required to bring portions of the mineral resource into mineral reserves, once a mine plan is in place and a pre-feasibility study completed.

Atna’s re-assaying program on the existing pulps from earlier 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 BSi Inspectorate of Reno, Nevada was performed using industry-standard analytical procedures, incorporating blind internal check samples and analytical standards 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 Xxxxxx Property.

The analytical database utilized to produce the resource model contained within this Technical Report was audited with minor error rates. Over 95% of the existing drill-hole intercepts were re-assayed by ALS Chemex with the re-assay results within acceptable ranges.

Methods utilized to determine the resources models for the Range Front and CX mineral zones at the Xxxxxx Project by Xxxxxx Xxx are consistent with practices in standard use in the industry.

1.7   Recommendations

The next exploration phase at the project should include the development of underground access to the Range Front and CX mineral zones. Development of the underground access will provide platforms for detailed underground drilling of the both the CX and Range Front zones. An initial drill spacing of 100-feet by 100-feet is recommended to be completed in the shallow portion of the Range Front zone to refine the understanding of the geology, structural controls and detailed distribution of high gold grades, and to obtain additional sample material for metallurgical testing. To be most cost-effective, the drilling should be carried out from both the surface (drill holes less than 1,000 feet deep) and from underground access.

Development of the underground access will also provide opportunities to define mining and development ground conditions and therefore refine the economic model and mine plan for extraction of the mineral resources at the property. Underground access will also allow the bulk sampling of the CX and Range Front zones and underground characterization of the mineral zones and evaluation of the mining methods within each of the zones.

All underground workings should be mapped in detail (1:240). A face and rib sampling program should be maintained throughout the underground drifting program and the collected data added to the resource modeling database.

If not accomplished via the underground definition drilling program, a series of six (6) reverse circulation rotary (RC) holes should be selected and “twinned” with diamond drill holes. This validation process, if the core results are similar to the original RC samples, will provide confirmation of the RC sample results. Each of these twin holes should be within 10 feet of the intercepts in the original RC drill holes.

Detailed surface mapping and sampling of the CX mineral zone is recommended within the CX pit (where pit wall and bench conditions are safe). Mapping at a scale of 1:240 and continuous channel sampling of the bench walls are recommended. Results from this work can then be added to the resource database and models.

Because drilling is rather widespread in the deeper portions of the Range Front and CX zones, it is recommended that the drill-hole density be increased to 200-foot centers. Completing this work will require a large amount of surface drilling with drill holes averaging approximately 1,300 feet in length. Alternatively, these holes can be drilled as underground access is developed to increase the drill density of the two zones and be completed as a development decline is completed.

Additionally, drilling along strike and down the dip of the zones is warranted as both the CX and Range Front zones remain open to expansion.

The costs associated with the recommended additional work at Xxxxxx are summarized as follows:

Surface and Underground diamond drilling (32,000ft)	US$1,622,500
Metallurgical Testing	US$100,000
Underground development (3,500ft drift)	US$4,500,000
Dewatering (2 xxxxx)	US$2,500,000
Permitting	US$650,000
Project Team - personnel	US$650,000
Land costs	US$250,000
Miscellaneous costs	US$400,000
Contingency @ ~12%	US$1,324,500
Proposed Project Budget	US$12,000,000

2.0 Introduction and Terms of Reference

2.1 Introduction

Xxxxxx Xxx, P.Xxx., is an independent consulting geologist engaged by Atna Resources to produce a mineral resource estimation for the Range Front and CX Zones located within the Xxxxxx gold property. Xxxxxx Xxx, a licensed Professional Geoscientist in the Province of British Columbia, Canada, is a Qualified Person and is responsible for the preparation of the mineral resource estimate and the subsequent supervision of the preparation of this Technical Report.

Xxxxxx Xxx received assistance in the preparation of this report from several of Atna Resources personnel or contractors including the following:

·   Xxxx X. Xxxxxxx, Site (contract) Geologist: Database and Sample Protocol Management

·   Xxxxxx X. XxxXxxxxx, Site (contract) Geologist: Contributor

·   Xxxxx X. (Xxx) Xxxxxxx, CPG #5138: Site (contract) Geologist and Principal Editor

·   Xxxxxxx X. Xxxxxxx L.G., MBA, Vice-President, Exploration, Atna Resources. Xx. Xxxxxxx assisted in the preparation of this report which is not directly related to the generation of the resource block model or recommendations and conclusions.

This report is based on drilling and sampling (assay) information as of February 15, 2005. This report is based on information known to Atna Resources Inc. as of March 23, 2005. All measurement units used are in Imperial (feet, ounces gold per short ton, tons) and any reference to currency is expressed in United States dollars.

2.2 Terms of Reference

Xxxxxx Xxx is not associated with Atna Resources, or any associated company. Xxx was paid a fee for his work on this Technical Report and his evaluation and development of the mineral resource model that is the focus of this Technical Report. The fees paid Sim for this work were at the rates that are customary and normal in the industry for an engagement of this sort and for work of this type.

Xxxxxx Xxx, in preparing this report, relied upon the extensive geologic and analytical database developed by previous operators (both exploration and production data). The sources of information are listed in the Reference section at the conclusion of this Technical Report and are for the most part housed at the Xxxxxx Mine site in the project archives. Xxxxxxx Xxxxxxx provided on-site, periodic oversight of the project and its personnel throughout the first phase of work by Xxxx, and Xxx visited the site September 22 through 25, 2004. Sim reviewed the QA/QC results and program for adequacy.

Xx. Xxxxxxx, Xx. XxxXxxxxx, and Xx. Xxxxxxx are independent consulting exploration and mining geologists with +14, +25, and +30 years experience respectively in the mineral industry.

3.0 Disclaimer

The results and opinions expressed in this report are based upon the authors’ field observations, the geological and technical data listed in the References and general technical expertise in the field of geology and mineral deposits. Xxxxxx Xxx has carefully reviewed the information provided to him by Xxxx and its consultants and believed the information used by him to develop the mineral resource model to be reliable.

Sim reserves the right, but will not be obliged, to revise the conclusions and recommendations made in this report, particularly with respect to the development of the resource model and the results of the mineral resource calculations, if additional information becomes available and know to the author subsequent to the date of this report.

The opinions and results reported within this Technical Report are conditional on available geologic and analytical data being current, accurate, and complete to the effective date (February 15, 2005) of this report.

4.0 Property Description and Location

4.1 Location

The Xxxxxx Mine site is located 27 air miles (43 kilometers) northeast of Winnemucca, in southeastern Humboldt County, Nevada, on the east flank of the Xxxxxx Mountains (Figure 4-1). Additionally, the property is located in approximately 35 driving miles east-northeast of the town of Winnemucca, Nevada, or 50 driving miles west-northwest of the town of Battle Mountain, Nevada. Winnemucca itself lies 175 miles northeast of Reno, Nevada.

The Xxxxxx properties (Xxxxxx Mine, Xxxxxx Mine, and Xxxxxx Xxxx) owned by Xxxxxx Mining Company (PMC) lie along the eastern flank of the Xxxxxx Mountains in the Potosi and Golconda mining districts of eastern Humboldt County, Nevada representing a key position on the Xxxxxxxx trend. Approximately 34,100 acres of fee land, claims and leases make up this land package. Atna Resources has rights to the northern portion of PMC’s land position centered at the Xxxxxx Mine and controls approximately 12,765 acres through its agreement with PMC.

Figure 4-1. Location Map of the Xxxxxx Mine Project

(Source: Atna Resources, March 2005)

4.2 Land Status and Agreements

The Xxxxxx property is made up of a number of property parcels that are either wholly owned by Xxxxxx Mining Company or under lease/option by Xxxxxx Mining Company and therefore subject to the Atna-Xxxxxx Mining Company (“PMC”) earn-in agreement. The property includes 3,800 acres of patented fee lands wholly owned by PMC, 360 acres of leased patented fee lands, 8,496 acres of federal unpatented lode mining claims wholly owned by PMC, 1,362 acres of leased federal unpatented lode claims. A total of 553 unpatented federal lode mining claims (both owned and leased by PMC) are included in the property position. Total acreage controlled by Xxxxxx Mining Company and subject to Atna’s earn-in agreement is 14,018 acres.

Figure 4-2, Land Status, shows the general property positions held by PMC and subject to the earn-in agreement between Atna Resources and Xxxxxx Mining Company.

Figure 4-2. Generalized Land Status Map

(Source: Atna Resources, Inc., March 2005)

A text summary of the land position is shown in Appendix 4-1, Xxxxxx 2005 Land Budget Projection, and in extended table form in Appendix 4-2, Exhibit A Part I, Unpatented Mining Claims and Appendix 4-3, Exhibit A Part II, Fee Lands.

The Company has received a title opinion on the core area of four square miles of the Xxxxxx property, within which all currently identified mineral resources and currently perceived exploration potential of the property exists, rendered by Xxxxxxx Xxxxxxxx of Xxxxxx & Xxxxxxxx, Reno, Nevada (See Figure 4-3 for Area of Detailed Title Opinion).

4.2.1 Xxxxxx Mining Company Agreement

Xxxx and PMC executed an Exploration Agreement with Option for Mining Venture on August 12th, 2004. The agreement allows Atna, through exploration and property maintenance expenditures, to earn up to a one hundred percent (100%) interest in the project over a four (4) year period. Under the terms of the agreement, Atna is obligated to spend US$1.5 million in the first year of the agreement on exploration and property maintenance on the Xxxxxx Property (this phase of the earn-in was completed by Atna prior to December 31st, 2004). Additionally, Atna must spend an additional US$10.5 million during the next three years of the agreement (US$12.0 million total) and deliver to Xxxxxx Mining Company a pre-feasibility study on the mineral resources evaluated on the property. Once the US$12.0 million in exploration expenditures are completed and the pre-feasibility study is provided to PMC, Atna will vest in a 70% joint venture interest in the property. Figures 4-2, Generalized Land Status and Figure 4-3, Detailed Land Status and Agreement Map, display the Area of Interest for the Exploration Agreement with Option for Mining Venture and the land holdings within the Area of Interest.

After Xxxx’s completion of the required earn-in expenditures and delivery of the pre-feasibility study to PMC, the agreement allows PMC several options going forward. PMC may elect to:

1) Participate as a 30% joint venture partner in the continued development of the project (Atna remains operator);

2) Back-in to a 70% joint venture interest in the project via the expenditure of US$30.0 million over a 36 month period (PMC would be operator during the back-in period); or

3) Elect to offer to sell its 30% interest in the project to Atna for US$15.0 million. PMC must make its election within 60 days of Atna’s completion of its earn-in requirements. If Barrick elects to sell its interest to Atna, Atna may either elect to purchase the interest or to not purchase the interest. If Atna elects to purchase PMC’s remaining 30% interest, Atna has a nine-month period to arrange the financing to close the purchase.

Figure 4-3, Detailed Land Status and Agreement Map

(Source: Atna Resources, Inc. & Xxxxxx Mining Company, August 2004 - March 2005)

4.2.2 Underlying Agreements - Unpatented Federal Mining Claims

Several lease agreements are in place to control portions of the properties comprising the Xxxxxx Project land holdings. Because of the complex ownership and royalty interests in the project area, all readers should consult Figure 4-3, Detailed Land Status while reading the summary text of the underlying agreements in Sections 4.2.2, 4.2.3, and 4.2.4. Additionally, Table 4-1, Land and Royalty Summary, follows Section 4.2.4 and summarizes the complex land status and royalty agreements related to the property subject to the Atna-Xxxxxx Mining Company agreement.

4.2.2.1 BEE DEE group

Newmont Capital Limited (50%) and X. X. Xxxxxx Trust (50%) are the underlying owners of the BEE DEE group of 58 unpatented federal mining claims located in Sections 4, 5 and 6 of Township 37 North, Range 42 East and Sections 31, 32, and 33 of Township 38 North, Range 42 East. The agreement provides for annual minimum advanced royalty payments, paid in monthly installments, to the owners totaling US$73,608 and the maintenance of the unpatented claims. Maintenance fees to the Bureau of Land Management and Humboldt County are current and the next fees, US$7,476.00, will be due in August and October 2005. The agreement expires March 8, 2010 unless extended by either PMC or Atna by written notice for an additional 10-year term.

The agreement calls for a royalty of 4% Net Smelter Return until royalties totaling US$500,000 are paid to the owners. Thereafter, the royalty is 2% Net Smelter Return. All advanced production royalty payments are credited against actual production payments due.

4.2.2.2 X. X. Xxxxx Unpatented Claims

Xx. Xxxxx is the underlying owner of 23 unpatented lode claims held by PMC under the terms and conditions of a Mining Lease agreement which expires on February 14, 2011 unless renewed by PMC by written notice. Of the 23 claims covered under the lease, 5 claims are within the Xxxxxx Project area of interest and are part of Atna’s property (Section 20, Township 38 North, Range 42 East). The terms of the agreement call for an annual advanced royalty payment of US$506 (pro-rata share of claims within the Xxxxxx Project area of interest) per year and PMC is required to maintain the unpatented claims in good standing. Maintenance fees to the BLM and Humboldt County are currently US$3,071 per year, with the next fees due in August and October of 2005.

The lease carries a mineral production royalty of 5% Net Smelter Return (conventional milling) or a 4% Net Smelter Return (heap xxxxx) royalty. All advanced production royalty payments are credited against actual production payments due.

4.2.2.3 VEK/Xxxxxx Unpatented Claims

VEK Associates and Xxxxxx Resources Corporation are owners of 56 unpatented lode claims under lease by PMC and subject to the Atna-PMC agreement. Term of the lease continues through June 14, 2008 unless extended by written notice by PMC. The lode claims are located in Sections 4, 8, and 16, Township 37 North, Range 42 East. PMC’s lease agreement with VEK/Xxxxxx includes provisions for a minimum annual advanced royalty payment of US$25,000 (currently payment is US$31,000), indexed to the consumer price index. Maintenance of the unpatented mining claims is the responsibility of PMC and will be US$7,076.00, due in August and September, 2005.

The agreement calls for a production royalty of 4% Net Smelter Return on all mineral production from the subject property. All advanced production royalty payments are credited against actual production payments due.

4.2.3 Underlying Agreements - Patented Fee Lands

Three lease agreements are in place controlling three separate patented fee land parcels within the Atna-PMC area of mutual interest at the Xxxxxx Property.

4.2.3.1 Xxxxxxxxxx, et al Lease

A 58.3333% interest in a 120 acre parcel in the southwest quarter of Section 28, Township 38 North, Range 42 East, is subject to a Mining Lease agreement between PMC and X. Xxxxxxxxxx (50% interest) and X. Xxxxxx (8.3333% interest). The other 41.6667% interest in this patented land parcel is owned directly by PMC (subject to a 2% retained Net Smelter Return Royalty by the sellers on their pro-rata undivided interest). The Xxxxxxxxxx lease provides for annual advanced royalty payments totaling US$17,172 per year (payable monthly on a pro-rata ownership basis).

The agreement also provides for a Net Smelter Return Royalty of 3% until US$1,500,000 in royalty payments have been made upon which the production royalty rate decreases to 2%. All advanced royalty payments are credited against any future actual production royalties.

4.2.3.2 X. Xxxxx Lease

The Xxxxx lease covers a 40 acre patented fee land parcel described as the southeast quarter of the northeast quarter of Section 13, Township 37 North, Range 42 East. The lease calls for an advanced minimum royalty payment of US$1,000 annually from January 1, 2006 through January 1, 2010 and US$1,500 annually from January 1, 2011 through January 1, 2015. Unless extended by mutual agreement, the lease will terminate on December 31, 2015.

Included in the agreement is a retained production royalty of 5% Net Smelter Return for milled ores and 4% of Net Smelter Return for heap leached ores or other lower recovery processing, payable to Xxxxx. No mineral production has occurred on the property to date.

4.2.3.3 Xxxxxx Lease

PMC leases 1,160 acres of land from X. Xxxxxx in the region and 320 acres of the lease are a part of the Xxxxxx Project lands. The lease cost per year to the Pinson Project is US$3,040, and is recoupable from actual production royalty payments for those payments made after January 1, 2002 ( payments are paid annually). The lands subject to this agreement are more specifically described as the south one-half of Section 13, Township 37 North, Range 42 East. The agreement expires on December 31, 2013 unless extended by written notice to Xxxxxx.

The lease calls for a 5% Net Smelter Return production royalty on all minerals produced from the property. To date, no mineral production has been made from the Xxxxxx portion of the property.

4.2.4 Underlying Agreements - Other Agreements

Several other agreements affect the lands subject included in the Atna-PMC agreement including the Xxxxxxxxxx Seven Dot Range Management agreement.

4.2.4.1 Xxxxxxxxxx - Seven Dot Agreement

This agreement provides annually US$10,000 to be deposited and drawn upon for range improvements by the Xxxxxxxxxx’x Seven Dot cattle operation. Only 50% of this cost is borne by the Xxxxxx Project with the remainder paid by PMC for the remaining acreage which includes lands near its Xxxxxx project to the south of the Xxxxxx property. This fund was set up to allow PMC to assist and mitigate effects of its operations on the Xxxxxxxxxx cattle business and avoid the continual requests for assistance from its neighbors, the Christisons, to provide backhoes, grader, or bulldozer work to benefit their cattle grazing business.

4.2.4.2 Xxxxxxx (Glamis Gold) Acquisition

Xxxxxxx Gold and Homestake Mining (co-owners of Xxxxxx Mining Company at the time) purchased Xxxxxxx’s (now Glamis Gold) interest in Xxxxxx Mining Company on November 30, 1996 and Homestake became operator of the Xxxxxx Mine (which was in production at the time). Subsequent to this purchase, Xxxxxxx Gold purchased Homestake and therefore PMC is not a wholly owned subsidiary of Xxxxxxx Gold.

This purchase transaction included an overriding royalty interest in the lands controlled by Xxxxxx Mining Company at the time of the transaction. This overriding royalty varies dependent upon the type of land (patented fee versus unpatented claims) and any underlying royalties existing on both owned and leased parcels at the time of the transaction. For fee lands owned by Xxxxxx Mining Company, Rayrock would be paid a 2.5% Net Smelter Return Royalty on lands not subject to an underlying royalty and 0.5% on those parcels subject to a royalty which would increase to a 1% Net Smelter Return royalty if the average gross value per ton of ore produced was greater than US$175/ton.

On fee lands leased by Xxxxxx Mining Company and subject to royalties payable to a third-party, the Rayrock royalty would vary from a minimum of 0.5% Net Smelter Return to a maximum of 5% Net Smelter Return as a result of the underlying royalty. The royalty percentage to Rayrock will be determined as the difference between a total royalty load of 6% less the underlying royalty; however the royalty will not exceed 5% or be reduced to less than 0.5% Net Smelter Return. For example, if the underlying royalty is 4%, then the Rayrock overriding royalty would be 6% less 4%, or a 2% royalty payable to Rayrock. If the underlying royalty is 0.5%, the Rayrock royalty would be 6% less 0.5% = 5.5%, which is greater than 5%, which would then be reduced to 5%. If the underlying royalty is 6% or greater, the Rayrock royalty interest would be limited to 0.5% of the Net Smelter Return.

If the lands were controlled by unpatented lode mining claims and not subject to underlying third-party agreements with retained royalties, the retained production royalty would be 2.5% Net Smelter Return. If the unpatented mining claims have underlying retained royalties, then the Net Smelter Return royalty would be determined as is the royalty described above under the patented lands leased by PMC and subject to an underlying royalty, again with a maximum of 5% Net Smelter Return and a minimum of 0.5% Net Smelter Return royalties dependent upon the underlying royalty load.

Additionally, there is a production royalty “holiday” from production from existing pits and extensions of mineralization within the pits applicable to this agreement. The initial “holiday” was limited to 200,000 ounces of gold production and there remains approximately 80,000 “holiday” gold ounces remaining prior to the project incurring any liability on this overriding royalty.

4.2.4.3 Cordilleran Overriding Royalty

Cordilleran Exploration, the original developers of the Xxxxxx Project, hold an overriding royalty position on several parcels including the patented lands of Sections 21 and 29, Township 38 North, Range 42 East. Cordilleran has a 3% Net Smelter Return production royalty on Section 21 and a 5% Net Smelter Return royalty on Section 29. Cordilleran also holds an overriding royalty on the Cord and Pacific unpatented lode mining claims of 5% of Net Smelter Return. Both these parcels are in turn subject to the Rayrock overriding royalty. The Cordilleran royalty also covers the Tip Top claims in the south half of Section 30, Township 38 North, Range 42 East, and the Pacific claims making up the majority of the west half of the northwest quarter and the northwest quarter of the southwest quarter of Section 28, Township 38 North, Range 42 East.

4.2.4.4 Xxxxxxxx Gold/Newmont Capital Overriding Royalty

PMC purchased three and three-quarter square miles of fee lands from the Xxxxxxxx Mine. Xxxxxxxx retained a 5% Net Smelter Return Royalty on these parcels (Sections 23, 27, 33, and the west one half and northeast quarter of Section 25, Township 38 North, Range 42 East. Additionally, Newmont Capital controls (via its acquisition of Euro Nevada Corporation) an additional 2% Net Smelter Return royalty on these parcels.

4.2.4.5 Echo Bay Lands

PMC purchased unpatented lode mining claims owned by Echo Bay Mines (now Kinross) in 1998. Echo Bay retained a 2.5% Net Smelter Return Royalty on the claims sold to PMC. Additionally, these lands were subject to a finder’s fee agreement between Echo Bay and Xxxx Xxx where Mr. Xxx is due US$5,000 per year or a 1% Net Smelter Return Royalty for bringing the claims to Echo Bay. Mr. Xxx’s royalty interest is subject to a maximum payment of US$135,000 in royalties or advanced royalties. Mr. Xxx is due a minimum annual payment of US$5,000. Lands subject to this royalty override are the claims located in the Sections 24, 26, 35 and the north half of Section 36, Township 38 North and Section 30, Township 38 North, Range 43 East.

4.3 Royalty Summary

The Xxxxxx property package is made up of 38 property parcels and nearly all are burdened by one or more retained mineral production royalties. Each parcel’s royalty agreement (or agreements) is structured slightly different than others and therefore the royalty for any given parcel is likely to be slightly different than an adjacent parcel. The mineral resource, the primary focus of this technical report, occurs entirely within Sections 28, 29, 32, and 33; Township 38 North; Range 42 East. Due to the relatively high-level of uncertainty for resource-level tonnage and grade calculations, particularly when the majority of the resource is categorized as “Inferred”, Atna has not yet broken down the resource into individual property parcels. Additionally, many of the retained production royalty agreements include advanced royalty payments (see above sections) that are recoupable from actual production royalty payments, therefore initially the actual royalties paid upon commencement of production from any given parcel may be significantly less than the stated royalty rate defined in the table below. Ultimately, should a minable reserve be established, Atna will determine exactly what resource blocks are located on which parcels such that an accurate mine plan, royalty payment burden, and cash flow model for the project may be established as part of a full feasibility study on the project’s mineral resources and or reserves (if reserves are established in the future).

Table 4-1 is a summary of the individual royalty burden on all parcels subject to the Atna-PMC earn-in agreement:

Table 4-1: Royalty Summary

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
	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
		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
	5-lode claims-PMC	95	5.0% NSR
Section 32 (see footnote 1)	18-lode claims	370	2.0% NSR
	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
	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
	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
	18-lode claims	280	2.5% NSR 1.0% NSR-capped at $135,000
	Total Acreage	14,018

Table Footnotes

1)   The resource announced in February 2005 by Atna Resources, Ltd. for the Xxxxxx project and supported by the Xxxxxx Gold Property Technical Report (NI 43-101, Technical Report filed in March 2005 with Sedar) is located with in four (4) square miles highlighted in the land/royalty table above (Sections 28, 29, 32, and 33 of Township 38 North, Range 42 East). Due to the relatively high-level of uncertainty for resource-level tonnage and grade calculations, particularly when the majority of the resource is categorized as “Inferred”, Atna has not yet broken down the resource into individual property parcels. Ultimately, should a minable reserve be established, Atna will determine exactly what resource blocks are located on which parcels such that an accurate mine plan, royalty payment burden, and cash flow model for the project may be established as part of a full feasibility study on the project’s mineral resources and or reserves (if reserves are established in the future).

5.0 Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1 Accessibility

Road access to the Xxxxxx Mine is by traveling east from Winnemucca on Interstate 80 for 15 miles (24 km), turning north at the Golconda exit, proceeding through Golconda to Nevada State Highway 789 and continuing to the end of the paved road (16 miles). At that point the road forks and is unpaved. The right fork leads to the Xxx Xxxxxx Mine and the townsite of Midas. The left fork leads to Xxxxxx Mine (4 miles), Xxxxxxxx and Turquoise Ridge Mines (11 miles) and terminates at Twin Creeks Mine (15 miles).

5.2 Climate

There are 11 climatologic stations within a radius of approximately 80 miles from the Xxxxxx Mine project area. These stations include: Battle Mountain, Xxxxxx Newmont Mine, Golconda, Gold Banks Mine, Xxxxx, Xxxxxxx Creek, Paradise Valley Ranch, Paris Ranch, Rye Patch Dam, Sleeper Mine and Winnemucca, the closest weather station to Xxxxxx Mine. Climatological records from these stations range in duration from 2 to 65 years. Precipitation data from the Xxxxxx Mine are also available from 1980 up to the present. Based on the available data, the average monthly and annual precipitation, temperature, and evaporation for the Xxxxxx Mine project area were developed. (Water Management Consultants, 1998).

Precipitation, although low, occurs steadily from October through June, and falls significantly in July through September. In a typical year, there are 27 days with measurable precipitation, amounting to 16” of snow and 8” of rain.

Based on 30-year averages, these are the expected seasonal variations in temperatures:

Table 5-1: Seasonal Temperature Variations

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)

5.3 Local Resources and Infrastructure

5.3.1 Background

Nevada is a mining state and well accustomed to mining. Humboldt County itself is entirely rural, with one population center, Winnemucca, the county seat, located in the southeastern part of the county. Winnemucca describes itself as a "high desert city full of classic charm," which has been continuously occupied since about 1830 when it became a base camp for trappers, and later as a way-station for pioneers migrating to the California goldfields after 1849 (Winnemucca, NV web site, 2005). Historically, Winnemucca was a local ranching community which branched out into support for large-scale mining following the discovery of significant gold deposits in the 1980s. Today, Winnemucca continues in that role, and has evolved as well into a jumping-off point for local outdoor recreation, as it is surrounded by vast tracts of Federally-administered public lands.

5.3.2 Physical Infrastructure

The major highways are paved or well maintained. Interstate 80 traverses the northern third of Nevada, from Wendover, Utah to Verdi, California. State Highway 789, leading from Golconda northeast towards Xxxxxx, is paved to within 4 miles of Xxxxxx. The unpaved portion, which begins as a "Y" junction at the end of pavement, is considered a private road under the control of Newmont Mining Corporation. That road is well maintained and serves as the main access roads to both Newmont’s Twin Creeks Mine complex and Placer Dome’s Xxxxxxxx and Turquoise Ridge mines.

Winnemucca is served by a general aviation airport and daily passenger rail service. Bus service was terminated several years ago.

Landline dialup telephone service for the Internet is slow from the minesite (28.8 kb via modem, or less), but high-speed broadband is available through satellite. Cellular phone service is available, but contingent on the strength of receiving antennas, topography and lines of sight.

5.4 Adjacent Operations

Two active mine sites exist a few miles further north of Xxxxxx: the Xxxxxxxx/Turquoise Ridge complex at 5 miles, and the Twin Creeks complex at 10 miles. Both sites create much passenger vehicle, shuttle bus and freight-truck traffic. In particular, Newmont ships refractory gold ores from other Northern Nevada properties to Twin Creeks for processing, and all Xxxxxxxx/Turquoise Ridge ores are processed at the Twin Creeks facilities.

5.5 Physiography, Flora, and Fauna

5.5.1 Physiography

The Xxxxxx Mine is situated in the Great Basin physiographic province and the Basin and Range tectonic province. North-south mountain ranges and parallel intermontane basins characterize the area. The entire region is a closed drainage system with all the permanent streams flowing to interior "sinks" such as the Xxxxxx and Humboldt sinks, or interior lakes such as Pyramid and Xxxxxx. Elevations in the area range from about 4,000 feet above mean sea level in the basinal areas, to over 9,000 feet in the surrounding ranges. The terrain is generally moderate.

5.5.2 Flora

The Xxxxxx Mine area is located in a sagebrush vegetation community, with mixed sagebrush-grass zones occurring at higher elevations or in broad valleys. Sagebrush and shrub species observed in the compilation of earlier environmental assessments (EAs) for Xxxxxx include two varieties of sagebrush (big and low); three of rabbitbrush (rubber, green and low); bitterbrush; little leaf horsebrush; and desert peach. Grasses recognized include Xxxxxxxx bluegrass, cheatgrass, Basin wild rye, wheatgrass, needlegrass, pepperweed, Russian thistle, halogeton, phlox, lupine, balsamroot and Indian paintbrush (BLM, 2001).

5.5.3 Fauna

The area supports a large number of small mammals, in particular Xxxxxxx’x cottontail rabbit and the black-tailed jackrabbit. These are by far the most commonly observed animals. Other small mammals known in the area include the deer mouse, western harvest mouse, Great Basin pocket mouse, long-tailed vole, chisel-toothed kangaroo rat, least chipmunk, and Xxxxxxxx’x ground squirrel (BLM, 2001). Mule deer and antelope also inhabit the foothills. Predator mammals identified as occasional visitors would include mountain lion, bobcat, coyote and badger. Scorpions and bull snakes have been seen on the property (La Peire, personal communication, 2005)

The vegetation is sufficiently dense to provide habitat suitable for small ground- and shrub-nesting birds, such as horned larks, sparrows, sage thrashers, nighthawks, barn swallows, meadowlarks, magpies and flycatchers. Game birds include mourning doves, California and mountain quail, sage grouse, Hungarian xxxxxxxxx and chukars. Raptor nests are not in evidence, but it is likely that golden eagles, western burrowing owl, ferruginous hawks and prairie falcons are the principal avian predators (BLM, 2001). A pair of ravens has occupied the pit areas throughout Atna’s exploration program.

6.0 History

6.1 Summary

Mineral exploration and discovery developed fitfully in this valley, beginning with the post-Civil War discovery of gold and silver in the Xxxxxx Range, resulting in the organization of the Fremont Mining District in 1874 and its reorganization into the Potosi Mining District in 1878. The district as now defined extends from Xxxxx Creek Valley on the east to the south end of the Xxxxxx on the west, approximately diagonally southwest from T. 39 N., R. 42 W. to T. 37 N., R. 40 W. (Xxxxxxx, 1998, p. 57).

6.2 Initial Discovery

Within the valley, mineral production was not significant until the discovery of the Xxxxxxxx Mine in 1934. That discovery initiated a period of gold production there which continued intermittently until 1967. During this period local ranchers, Xxxxxxx Xxxx and Xxxxxx Xxxxxx, prospected and discovered small siliceous outcrops resembling Xxxxxxxx ore, and leased them to the Xxxxxxxx Mine owners, resulting in mining of approximately 100,000 tons in 1949-1950 from an initial pit exposing ore grading 0.14 ounce per ton (opt) gold (4.8 g/t) (Xxxxxxxxxx, 1983; XxXxxxxxx et al., 2000).

6.3 Pre-1970 Development

Exploration efforts between 1950 and 1970 were limited to two programs:

·   A 1963 drilling campaign by Nevada Goldfields Corporation to prove up a resource of 47,000 tons @ 0.17 opt (5.8 g/t)

·   A 1968 JV between Getty Oil and Homestake Mining to explore the Xxxxxx claims and complete 2 core holes

6.4 Cordex Syndicate Exploration and Development

The property remained functionally dormant until 1970, when an exploration group known as the Cordex I Syndicate (Xxxx Xxxxxxxxx, Xxxxx Xxxxx and Rayrock Resources) leased the property on the strength of its similarity to Xxxxxxxx and structural position along the range-front fault zone bordering the Osgoods. Following a surface mapping and sampling program, 17 rotary holes were completed in and around the initial Xxxxxx pit in 1971 and confirmed only low-grade gold values. An 18th step-out hole discovered a 90' (27.4 m) intercept of 0.17 (5.8 g/t), indicative of a subcropping extension to known mineralization northeast of the original pit, resulting in the definition of what would become the "A" zone at the Xxxxxx property - a 60 foot x 1,000 foot shear zone estimated to contain 1.5 M tons of 0.18 opt Au.

Continued exploration southwest of the original pit defined the "B" zone, estimated at 1.7 M tons at 0.15 opt Au. No production was attempted at the time because of low gold prices ($65 per oz) (Xxxxxxxxxx, 1983; XxXxxxxxx et al., 2000).

The gradual rise of gold prices to > $250 per oz. in the late 1970s allowed the Cordex I Syndicate - which then consisted of several minority partners - to reorganize into a partnership known as Xxxxxx Mining Company, with Rayrock Mines as the operator, and begin production. Homestake Mining Company and Xxxxxxx Gold became partners in Xxxxxx Mining through their purchase of some of the minority interests (XxXxxxxxx et al., 2000).

6.5 Xxxxxx Mining Company Exploration and Development

Xxxxxx Mining began development of the A Pit in 1980, and produced gold from it the next year. Production from the B Pit followed in 1982. Step-out drilling in 1982-1983 to the northeast of the A zone intersected 2 more discrete zones: the C zone extending ENE from the A zone, and the CX zone extending NE from the C zone. C zone production replaced the final B zone production in 1988, and CX production was delayed until 1990, owing to the slightly greater stripping requirement to reach ore.

Step-out drilling NE of the CX zone in 1984 located an apparently independent fault system (striking NNW, dipping steeply E) that became the core of the MAG Deposit, which went into production in 1987. In 1993, step-out drilling NW of the CX zone once again established more ore in proximity, a CX West zone subparallel to, and closer to, the Xxxxxx Mountain front. Xxxxxx Mining Company produced from the CX, CX West and MAG pits as long as possible in the mid to late 1990s, until a combination of falling gold prices and erratic mill feed forced closure of its oxide mill in early 1998. Continued attempts to expand production of oxide ore failed, and all active mining ceased on January 28, 1999 (XxXxxxxxx et al., 2000).

6.6 Homestake-Barrick Exploration

Homestake and Barrick conducted a JV exploration program from 1996 to 2000 expending some $12 million on the project. The JV then determined it would not expend any further funds on exploration. Subsequent to that decision, Barrick acquired Homestake Mining Company and drilled an additional three exploration drill holes in 2003.

6.7 Xxxxxx Project Production Summary

Historically, the district was credited with Au production in excess of 1 M oz, but less than 100,000 oz. Xx (Xxxxxxx, 1998, p. 108 & 120). Independently, Barrick compiled a similar record of production and credits the Xxxxxx mine property with production of 986,000 ounces of gold (Appendix 6-1, Discovery & Production History).

6.8 Atna Resources Exploration

Atna Resources began project planning during July of 2004 and began drilling at the Xxxxxx property in August 2004 after the execution of the earn-in agreement with Xxxxxx Mining Company on August 12^th.

7.0 Geological Setting

7.1 Regional Setting

For the last half-century, sediment-hosted, Xxxxxx-type gold systems have accounted for most economic gold discoveries in Nevada. Mineralization in these environments lies mainly in four geographic belts of mostly Paleozoic carbonate rocks. These belts are located in north-central Nevada, and the three most productive pass through the townsites of Carlin (“Xxxxxx Trend”), Battle Mountain (“Battle Mountain-Eureka Trend”) and Golconda (“Xxxxxxxx Trend”). The fourth belt, the “Independence Trend”, is located north of the town of Elko and is the location of the Jerritt Canyon group of mines (Queenstake Resources) and the Big Springs Mine (Golden Gate Resources). Collectively, these belts hold a geochemical endowment of over 200 million ounces of gold. A key to understanding this endowment lies in understanding the structural relationships of the stratigraphic section.

The Paleozoic era throughout north-central Nevada is a period in which a craton extended from the Mid-Continent westward to about the longitude of present-day Battle Mountain (Xxxxxxx, 1980). East of the inferred craton margin was a gently sloping (miogeoclinal) carbonate shelf, consisting of a variety of quiet-water quartzite, limestone, and intrabasinal shale (“Eastern Assemblage” rocks); and west of the margin was a much deeper oceanic environment depositing a sequence of siliciclastic sediments, largely composed of deep-water shale, chert, and volcanic ashes and flows (“Western Assemblage” rocks).

During the Paleozoic, sedimentary rocks were structurally deformed by a series of east-directed compressional tectonic events. There were at least two such events of primary importance: A Devonian-Mississippian event, transporting pre-Devonian deep-water sediments eastward ("Antler orogeny") created the Xxxxxxx Mountain and associated thrust faults. This was followed by a Permo-Triassic event, transporting Mississippian to Permian age deep-water sediments eastward ("Sonoma orogeny") creating the Golconda and Humboldt Thrust faults. A period of tectonic quiescence, allowing for the deposition of local basin sands and conglomerates and shallow limestones of the “overlap” sequence, marked erosion and subsidence during Mississippian to Permian times, and separated the two compressional events.

Both compressional events deformed rock units into a series of folds, and thrust the western assemblage siliciclastic rocks over the eastern assemblage carbonate rocks, forming one of the principal structural traps necessary for the formation of Xxxxxx-type sediment-hosted gold deposits.

Compressional tectonics continued until the earliest Tertiary with the emplacement of granitic stocks, the development of additional fold and thrust belts, and the onset of strike slip faulting. Much of this activity was west of the Battle Mountain-Eureka Trend, but extended as far east as Western Utah and Southern Wyoming.

Beginning in Tertiary time, about 48 Ma, extensional tectonics became the primary tectonic regime in the Great Basin. This structural style formed a series of horsts and grabens, resulting in the present-day landscape of northerly oriented parallel mountain ranges and valleys (Xxxxxxx, 1980). Owing to the large number of Xxxxxx-type deposits with ages between 35 and 48 Ma (Arehart et al., 2003), the onset of extension has been considered by some (Ilchik and Xxxxxx, 1997) to be an important factor in the genesis of these deposit types.

Figure 7-1. Regional Stratigraphy

(Source: Atna Resources, March 2005)

In a general sense, this geologic history of north-central Nevada is reflected somewhat in the geology of the Xxxxxx Mountains. The range, as mapped by Xxxx and Xxxxxxx (1964), and Xxxxxxxx and Xxxxx (1974) shows a variety of rock units through time (Figure 7-1, Regional Stratigraphy). These initial workers show the oldest Cambrian-aged siliclastics (Xxxxxx and Xxxxxx Mtn. quartzite) to be overlain by Ordovician carbonate rocks (Comus Formation), and folded into a broad, north plunging anticline. The west flank of this anticline has been overthrust by deep-water siliceous shale and cherts of the Ordovician Valmy Formation, while in the core of the range, and in scattered localities on the east side of the range, sands and conglomerates of the Battle Formation and limestones of the Xxxxxxx Formation lie unconformably on, or are in fault contact with, these older folded rocks (Figure 7-2, Regional Geology).

A second structural event has been mapped by these early workers along the northwest and southern flanks of the Xxxxxx Mountains. This event displaces Mississippian volcanics and Pennsylvanian shales, marking the Golconda and Humboldt thrusts. Before the onset of extension in the Tertiary, Cretaceous stocks consisting of diorite and granodiorite were intruded into the core of the range, and created a large, thermally-metamorphosed aureole with several tungsten skarns surrounding the stock. Capping the sequence are Tertiary volcanic rocks consisting of older rhyolite tuffs, Miocene basalt and andesite flows, and younger basalt flows.

Due to structural complexity and lithologic variations within the range, more recent workers (Xxxxx, 1991, Xxxxxx-XxXxxxx and Xxxxx, 1991, and XxXxxxxxx et al., 2000) have distinguished terranes based on age, structural history, and lithology. Four principal terranes have been identified (Figure 7-3, TectonoStratigraphy).

These are:

·   An Xxxxxx Mountain terrane, consisting of the structural block cored by pre-Cambrian/Cambrian Xxxxxx Mountain quartzite in the southern Osgoods, and in the Xxxxxx-area, Cambrian-Lower Ordovician Xxxxxx phyllites, Ordovician Lower and Upper Comus, and an intrusive Cretaceous granodiorite stock. The domain of these rocks extends northward, from Xxxxxx through Xxxxxx to Xxxxxxxx. Further northeast, at the Turquoise Ridge and Twin Creeks Mines, the Upper Comus is additionally overlain by a local assemblage (Twin Creeks Member) of calcareous shales intruded by numerous mafic xxxxx.

The remaining three terranes are structural overlap sequences which generally crop out further north:

·   A Leviathan allochthon terrane, presumably Ordovician, exposed only at Turquoise Ridge and Twin Creeks, characterized by cherty, basaltic, and pelagic (tuffaceous) sequences thrust over the Twin Creeks Member.

·   An "Antler overlap" sequence, consisting of a regional Battle Mountain quartzite-rich conglomerate and overlying Pennsylvanian-Permian Xxxxxxx calcareous sandstone/fossiliferous limestone pairing, all of which are locally absent at Xxxxxx.

·   A Golconda allochthon terrane, transporting Mississippian Goughs Canyon and Penn-Permian Farrel Canyon sediments, but restricted areally to the northern part of the Xxxxxx Mountains, far from Xxxxxx proper.

Figure 7-2 Regional Geology

(Source: Atna Resources, March 2005)

Figure 7-3. TectonoStratigraphy

(Source: Atna Resources, March 2005)

7.2 Xxxxxx Mine Setting

The Xxxxxx Mine site lies on the east flank of a large stock of Cretaceous granodiorite which cores the southern part of the Xxxxxx Mountains (Figure 7-4, Mine Site Geology). Sediments peripheral to the east side of the stock dip moderately to steeply southeast, east and northeast, depending on their circumferential position. The lowest stratigraphic units preserved against the granodiorite contact are Cambrian Xxxxxx phyllitic shales, limestone interbeds, and variously hornfelsed sediments. These are overlain by a thick sequence of variegated Ordovician Comus sediments with a significant carbonate content. Generally, the lowest parts of the Comus are massive, micritic limestones; the middle portion is dominated by mixed limestone shale interbedded sequences, often combined with local debris flows; and the upper Comus is characterized by mildly to non-calcareous shales.

The Comus Formation is in enigmatic depositional relationship to the underlying Xxxxxx Formation. It is routinely believed to be conformable or depositional, but in fact at Xxxxxx Mine the contact is for the most part faulted, subparallel to the envelope of the Range Front Fault.

The Cretaceous granodiorite stock, emplaced at 90-92 Ma (Xxxxx et al., 1997, cited in XxXxxxxxx et al., 2000) has created an irregular contact metamorphic aureole extending up to 10,000 feet (3,000 m) from its perimeter. The effects on the metamorphosed sedimentary section vary:

·   Much of the Xxxxxx section is hornfelsed;

·   The Upper Xxxxxx in particular is converted to a maroon, biotite-cordierite hornfels;

·   The Lower Comus is commonly calc-silicated (converted to wollastonite and idocrase);

·   The Upper Comus may be converted to chiastolite hornfels.

The most important structural feature of the Xxxxxx Mine area is the network of faults that borders the escarpment marking the southern and eastern edge of the Xxxxxx granodiorite. This fault system has been variously interpreted as a single master fault (Range Front Fault) curving around the stock or, more likely, a network of shorter, straighter segments that collectively accommodate several thousand feet of displacement while making a 50 degree bend around the SE corner of the stock (Xxxxxx Fault on south margin; CX Fault on southeast margin; Range Front Fault on east margin). Sedimentary rocks in the vicinity of this curvilnear system generally dip steeply away from the contacts of the granodiorite.

The relation of the Range Front and CX faults to Xxxxxx stratigraphy may be characterized as follows:

·   The "floor" of the complex is the Xxxxxx granodiorite;

·   The Cambrian Xxxxxx may be in intrusive contact with the granodiorite

·   The Range Front fault will mark the approximate location of the contact between the Cambrian Xxxxxx and Ordovician Comus. The contact zone will be brittlely broken or ductilely sheared.

·   The CX fault zone will separate Lower Comus from Middle Comus.

·   The Upper Comus may be folded in a synform between the Range Fault and CX faults, and be involved in some Range Front mineralization, but will not be geometrically involved in any CX faulting.

In addition to the large-scale fault system described above, there are numerous NW and a few E-W structures identified in past mapping and drilling. Generally, these appear to be mostly older than, and truncated by, the main system. Some of these faults have been reactivated enough to disrupt the continuity of the main Xxxxxx system.

North of the Xxxxxx Mine proper, a few core holes have intersected definite repetitions of the Xxxxxx/Comus stratigraphic section. It is probable that a sole thrust locally places Xxxxxx over Comus. Imbricate systems of this nature produce thrust duplexes which are highly susceptible to mineralization. Well-exposed examples are currently being mined in, for example, the Xxxxxx District south of Battle Mountain (Xxxx et al., 2004).

Figure 7-4. Mine Site Geology

(Source: Atna Resources, March 2005)

8.0 Deposit Type

8.1 Sediment-hosted, Xxxxxx-type Gold System

Gold mineralization at the Xxxxxx Mine is considered to be typical of ‘Xxxxxx-type,’ or sediment hosted gold deposits.

‘Xxxxxx-type’ mineralization was first recognized from discoveries in east-central Nevada in the early 1960s (Xxxxxx Mine). Similar deposits had been mined prior to the 1960s (for example Xxxxxxxx, first mined in the 1930s), but their discovery in Nevada in 1961, and the recognition of the gold endowment they represented, has led to the opening of more than 100 mines, producing about 200 million ounces of gold, assuring Nevada a leading position in world gold production (Xxxxx, 2004, p. 1).

Xxxxxx-type gold deposits display the following characteristics:

·   Gold occurs primarily as ionic substitution or micron-sized particles, often in xxxxxxxx pyrite, locally termed “sooty pyrite”.

·   Gold is hosted primarily by silty limestone to calcareous siltstone lithologies in the vicinity of major structural zones.

·   Gold mineralization is concentrated in structural traps and/or replacement horizons of receptive permeable beds.

·   Subtle alteration, dominated by decalcification and argillization of the host rock, and accompanied by selective silicification (jasperoid) and carbon-flooding.

·   “Gangue” (non-economic) minerals - calcite (calcium carbonate), siderite (iron carbonate), and-ferroan-dolomite (calcium-magnesium carbonate) - occur as geochemical fronts at many deposits, but are not ubiquitous.

·   “Pathfinder” elements (Sb, As, Hg) often occur in spatial association as the minerals orpiment, realgar, stibnite, or cinnabar.

·   Dikes, although not ubiquitous at some deposits, are characteristic at most (Xxxxxxxx, Goldstrike, Xxxxxx, Xxxxxxx Canyon) and occupy many of the mineralized fault zones in the deposits

Mass-balance analysis of plausible chemical reactions in individual deposits has led to a belief that (1) fluid-wallrock interactions and (2) sulfidation of reactive iron are important deposit-scale mechanisms of mineralization. Although this chemical feature is well-documented (Hofstra and Xxxxx, 2000; Xxxxxxx et al., 1998), the underlying reason for it is not.

Except for the gold-xxxxxxxx pyrite association, other features of Xxxxxx-type systems do not seem to be characteristic of a particular process (such as banded veins in epithermal systems) that would tie into a genetic explanation for the occurrence of the system (Seedorff and Xxxxxx, 2004). The deposits require sources of heat, gold, sulfur, and iron; a means of fluid transport; and receptive rocks. The receptive rocks, fortunately, exist over a large area of east-central Nevada, as explained earlier in the general review of Paleozoic stratigraphy. There are numerous occurrences of telescoped overthrust sections, both older-over-younger and younger-over-older, which place potential ore hosts in a variety of complex structural and stratigraphic settings, which create a variety of traps, and large regional structures cutting through these sections that serve as pathways for circulation of hydrothermal fluids.

One enigma remains the heat source. Both magmatic and metamorphic sources have been discounted, as there seems to be no direct age relationship to major stocks, and regional metamorphism would have likely occurred in the Late Cretaceous or Earliest Tertiary, when plate collisions, underthrusting of continental slabs and crustal-thickening events were active. Two theories are gaining currency, though:

1.   Regional Eocene magmatism was probably a major source of thermal energy throughout the Xxxxxx-type province (Xxxxxxxx and Xxxxxx, 2004);

2.   Basinal mechanisms, which rely on the onset of Basin-Range extension to create crustal permeability, could have facilitated the widespread circulation of heated meteoric water (Seedorff and Xxxxxx, 2004).

Ambiguous igneous and radiometric evidence exists to lend support to either of the above hypotheses: presumed Tertiary dikes occur at many of the deposits, and a great number of deposits have been dated at 39-42 Ma (Arehart et al., 2003).

9.0 Mineralization

Past production focused solely on oxidized ore. Oxidation is extensive in the CX-Xxxxxx fault system, affecting the entire length of the zone and penetrating up to 1500 feet in depth. The oxidation process generated pervasive limonite, hematite, and other iron and arsenic oxides. Oxidation is variably developed in the Range Front fault. In some fault and shear zones, oxidation is present at depths of 1800 feet; in other places oxidation gives way at shallow depths (<500 feet) to sulfide mineralization.

Sulfide mineralization is a pervasive event consisting of two stages of pyrite development, a non-ore pyrite, and an ore-stage xxxxxxxx pyrite. In hand specimens, gold bearing pyrite is a dull brassy to black color and extremely fine-grained. Remobilized carbon is usually associated with the pyrite, giving the sulfides a “sooty” appearance. Gold is primarily contained in pyrite, or found as rims around fine pyrite grains (Xxxxxxx and Wittkopp, 1983, Xxxxxx, 1994). Detailed geochemical work suggests there is also a primary positive Au-Hg (and weaker Au-As) correlation, as well as a negative Au-Ba association, which is preserved in the unoxidized environment (Xxxxxx and Xxxxxxxxxx, 1991). No other major elements (Pb, Zn, Cu, Mo, F) show any positive or negative relation to Au.

9.1 Host Rocks

The environments being targeted for continued exploration are shown in three representative cross sections which span the area of the current resource calculation and structures under study:

·   Section 6800 NE (Appendix 9-1) shows the principal relationships involving the CX fault and related strands.

·   Section 7100 NE (Appendix 9-2) is 300 feet further NE along strike of the CX fault zone, and picks up the definable beginnings of the Range Front fault zone.

·   Section 7700 NE (Appendix 9-3) is another 600 feet further NE along the same trend, and shows principally the Range Front fault zone.

In all three sections, several features of stratigraphy and structure are evident.

Host rocks for gold mineralization at Xxxxxx are a sequence of interbedded shale, siltstone, and limestone of the Upper and Lower Comus Formation of Ordovician age. These rocks are exposed in outcrop and the mine pits along the east flank of the Xxxxxx Mountains. The Upper Comus is a gently east-dipping complex consisting of black carbonaceous shale interbedded with varying, but minor, amounts of calcareous siltstone and silty limestone. Beneath the Upper Comus lies a large volume of rock assigned to the Lower Comus, which consists of a monotonous sequence of interbedded shale and limestone in roughly equal proportions. Local variations in the Upper and Lower Comus have been assigned unit names of “Ocls” and “Ocib” depending on relative abundance of limestone versus shale.

Within the Lower Comus interbedded shale/limestone unit, there are large irregular pods of Ordovician contact-metamorphosed carbonates, ranging from pure marble to garnet-epidote skarn to calc-silicated, wollastonite-dominated rocks. These are combined as skarn/marble (“Osm”) and shown separately on the cross-sections. Finally, the lowest defined stratigraphic unit is the Cambrian Xxxxxx, ordinarily a maroon colored hornfelsed phyllite in the Xxxxxx mine area.

Structurally, the cross-sections are dominated by two large fault systems. The CX fault zone, appearing on 6800 NE and 7100 NE, is a zone of brittle fracture within the Ocib package, and creates one dominant and several narrower splays of mineralization. The Range Front fault zone (displayed on cross-sections 7100 NE and 7700 NE), in contrast, is much thicker and persistent, and involves the Cambrian/Ordovician contact zone in extensive brecciation. On the sections, the contact zone is represented as a single line, but in reality is obliterated in many core intercepts by a broad zone of shearing.

Internally, the rocks display a significant degree of ductile deformation of undetermined age. Numerous instances of necking, rotation and dismembering of beds, polished slip surfaces, healed discontinuities and sheath folding are evidence of deformation under high confining pressures, and mask less extreme features, such as debris-flows and related soft-sediment deformation.

9.2 Alteration

Alteration, consisting of silicification, decalcification, argillization, introduced carbon and pyrite, is enhanced along structural zones. Primary alteration associated with the gold mineral system has been affected by a later stage of oxidation, with or without the development of clay fracture-fillings and veinlets.

The alteration suite at Xxxxxx displays features common to other Xxxxxx-type, sediment-hosted environments. Historical accounts of the Pinson Mine (XxXxxxxxx et al., 2000) describe alteration found along the CX and MAG faults.

In the CX environment, which includes the A, B, C, CX and CX West pits, there were gradational changes in the style and intensity of alteration. Beginning in the southwest, the B Pit was noted as a faulted interface between carbonates and argillites, with gold deposition occurring in weakly silicified and kaolinized fractures. Nearby, the A Pit was dominated by the development of gold-rich jasperoid (silicification of limestone-dominant rocks); within it, arsenical pyrite contained inclusions of <5 micron free gold. Accessory minerals were chalcedony, kaolinite, marcasite, pyrite and sericite. Further northeast, the C Pit hosted ore in decalcified carbonates cut by small cross-faults.

The CX Fault Zone, developed in the CX Pit, was characterized by a swarm of subparallel fault strands containing silica-pyrite replacements along narrow zones of carbonate. Silicification was more prevalent here than in the Range Front zone. A large volume of the adjacent hangingwall carbonate was thoroughly decalcified, but barren. Broad pervasive zones of alteration are not present within the CX mineral zone. Argillization and calcite veining occur in and adjacent to the fault zone. In many instances the fault breccia matrix has been completely argillized (Xxxxxxxxxx, 1983). A deep core test undertaken in 1993 (DDH-1541) encountered a leached, gold-bearing zone containing marcasite, arsenopyrite, and traces of cinnabar, realgar, stibnite, sphalerite, galena and native arsenic. The CX West Pit, a fault zone parallel to but 600 feet west of the CX, hosted ores extracted from the faulted contact zone separating lower, calc-silicated Comus carbonates from Upper Comus argillites.

The MAG pit developed near the projected intersection of the north-northwest trending MAG fault and the northeast trending CX fault. Certain lithologies (calcareous argillite and shale) were preferentially mineralized relative to others (pure argillite and calc-silicates). The dominant alteration elements were total decalcification and leaching of the carbonate host, accompanied by extensive argillization. Pervasive silicification healed fault xxxxx and breccias, produced jasperoids in the leached rock, and replaced undisturbed rock adjacent to the silicified structures.

The Range Front fault zone, in contrast to the CX-related system, displays primarily pervasive argillization and decalcification of the host rock. Where strongest, the zones consist of punky, spongy decalcified limestone in an argillically altered shaley, silty matrix. White and xxxxx xxxx fracture- and void-fill veinlets and patchy zones occur which give the rock a mottled appearance. Silicification appears minor and occurs as a broad overprint, perhaps related to diagenesis, or as altered clasts in breccia zones. Calcite veining is prevalent throughout the system but is strongest along the margins of the Range Front fault envelope.

In the Atna exploration program to date, mineralized core holes in both the CX and Range Front (RF) zones encountered the alteration features typical of past Xxxxxx mining experience, and expected in Xxxxxx-type systems, as detailed below:

Table 9-1: Inventory of Alteration Types in Atna Phase I Drilling

Alteration Type	Example CX Holes	Example RF Holes
Weak silicification	201, 204	208, 215, 223, 229A
Strong silicification	204, 206, 211, 213, 216, 219	209, 212, 221, 227, 230
Decalcification	206, 216	212
Argillization	203, 204, 211, 216, 220	202, 209, 212, 215
Carbon flooding	216, 219, 226	202, 221, 225
Secondary pyrite	219, 226	202, 207, 215, 221
Orpiment & realgar	226	202, 215, 223, 228
Oxidation	203, 219	230
Brecciation	204, 211, 216, 219, 220	205, 207, 209, 210, 212, 215, 221, 222, 223, 225

9.3 Mineralized Zone Configuration

Figures 9-1 and 9-2 show grade thickness contours defining the current known morphology of the CX and Range Front Fault mineralized zones in longitudinal section. The sections are oriented approximately N30E, with the left edge southwest, the right edge northeast, and the section dipping toward the viewer at 55 degrees. The sections represent the grade multiplied by the thickness of mineralization along the strike of the fault plane. Xxxxxx Mining Company drill holes are shown in green and recent Atna Resources drilling shown in yellow ochre.

The CX mineral zone (Figure 9-1) is the smaller of the two, with an approximate strike length of nearly 2000 feet and an overall down-dip extent in excess of 1500 feet, with grades averaging better than 0.25 oz/ton. From the section it can be seen that mineralization is poddy near the surface, with three separate zones of mineralization at and just below pit level. Below about 4500 feet in elevation, mineralization becomes more consistent, with two separate bodies in evidence. Both extend down dip over 600 feet, with strike lengths of 300 to 500 feet. These two zones are separated by a weak to non-mineralized zone. Causes for this barren zone are not fully understood, but post mineral faulting is a likely cause.

Atna Resources’ drill program in the CX was designed to define the extents of the southern mineral zone, and firm up the potential down dip extent in the northern mineral zone. Holes APCX-226, 224, 218 and 216 limited the extent of the southern mineral zone on the south and north edges, helping to refine the barren zone between the north and south zones. Hole APCX-219, arguably the best Atna drill hole into CX mineralization, added significant width and grade potential to the northern mineral zone. This hole also helps define a roughly 20 degree northeast oriented rake in thicker, higher grade mineralization. This thicker zone remains open at depth to the northeast.

Figure 9-1. Grade-Thickness Long Section, CX Fault Zone

(Source: Atna Resources, March 2005)

Mineralization along the Range Front Fault, in contrast to the CX Fault, appears to be more consistent (Figure 9-2). Drill spacing in the Range Front system is broad at this stage, but what appears to be developing is a near vertical zone of mineralization from about 4700 feet in elevation, down to approximately 3500 feet, with grades greater than 0.1 oz/ton gold. In the early stages of exploration, this columnar area appears to consist of two connected zones below 4400 feet, and a separate zone above 4500 feet. The area between the zones is weakly mineralized (see APRF-215), and may represent post-mineral offset along the CX West Fault, an ENE trending, west-dipping fault.

Prior to Atna’s program, mineralized intercepts in the Range Front zones were sparse, with holes spaced 400 to 600 feet apart. Atna’s program was designed primarily to fill in between these intercepts and step out from known thick high grade intercepts in holes HPR-050, HPC-162, HPC-144, and HPC-075 drilled by Homestake Mining Company (HMC) as operator of the Project for the Homestake/Barrick partnership. Atna’s program has begun to define an upper zone of mineralization with offsets of hole HPR-050 that carry mineralization up dip almost 300 feet to hole APRF-207, 400 feet to the southwest (hole APRF-209), and 200 feet southeast in hole APRF-212. Atna drilling at deeper elevations is also beginning to define a zone of mineralization surrounding holes APRF-225 and APRF-223. These two holes were drilled 400 feet downdip of hole HPR-050, halfway to intercepts at depth in HMC holes 144, 162, and 075. Hole APRF-221, targeted within the triangle defined by the last 3 holes, confirmed continuity of mineralization among these holes.

Figure 9-2. Grade-Thickness Long Section, Range Front Fault Zone

(Source: Atna Resources, March 2005)

As in the CX mineral zone, an apparent rake is developing in the thicker, higher grade intercepts which plunges NE from 10 to 30 degrees. The rake in both the CX and Range Front zones probably represents the trend and plunge of intersections between fault planes and favorable lithologic horizons.

10.0 Exploration

10.1 Introduction

Exploration approaches at Xxxxxx have consisted primarily of mapping, geochemical sampling and drilling. These approaches have been successful in discovering about 1 million ounces of gold in several exploited open pit deposits. Several geophysical approaches have been applied with limited or non-measurable success. Geophysics has largely been applied to exploratory programs along strike of known mineralized zones and as grass-roots applications to find additional mineralized zones.

10.2 Geologic Mapping and Geochemical Sampling

Cordex Syndicate, and its successor, Xxxxxx Mining Company, explored the property largely through mapping and geochemical sampling. There are three known mapping programs:

1) A regional mapping program from Xxxxxx to Xxxxxxxx by Xxxx Xxxxxxx in the late 1970s. The map has yet to be recovered from the Xxxxxx property archives.

2) A 1:6000 scale mapping program of the Xxxxxx property in 1983. Portions of the map series have been located in Xxxxxx property archives.

3) A 1:2400 scale mapping program of the Pit areas through the active life of the mine. Portions of the map series have been located in Xxxxxx property archives.

In addition, bench mapping in the pits occurred during mining, and was followed up by detailed 1:1200 scale maps of the A, B, C, CX, MAG, CXW, and Bluebell pits by Xxx Xxxxxxxx in 2000, after the cessation of mining activities. Xxxxxxxx’x maps were completed under the Homestake Mining/Barrick partnership agreement.

Several geochemical programs were completed by the Cordex Syndicate and Xxxxxx Mining Company during early discovery of the Xxxxxx Mine, throughout active mining, and by Homestake Mining Company. These are:

1) Cordex Syndicate took rock chip samples during ongoing mapping. A total of 737 samples were taken in this program. Samples were assayed consistently for gold, silver, arsenic, antimony, and mercury, with select samples also analyzed for lead, zinc, copper, and manganese. The dual programs of sampling and mapping were responsible for the discoveries of the Blue Bell and Xxxxx Canyon pit areas (Xxxxxxxx et al., 2000).

2) Pinson Mining completed 6 float chip geochemical grids consisting of 8,756 samples. These grids cover the MAG deposit, and on strike to the south of the A and B pits.

3) A biogeochemical study of sagebrush was run in the 1990s. Results from this study were inconclusive.

4) 312 additional rock samples and 273 additional soil samples were taken under the Homestake/Xxxxxxx XX program. These sample programs were completed on strike to the south of the existing pit areas, and to the west of the A, B, C, and CX pits, near the Granite Creek Tungsten Mine.

10.3 Drilling

There are several generations of drill programs on the property, beginning in 1963 with Nevada Goldfields Corporation’s drilling program.

1) Reverse circulation drilling by Nevada Goldfields Corporation in the 1960’s to prove up a small resource in the “B” pit area.

2) A 2-hole program (core) by Homestake Mining Company in 1968.

3) Exploration and development drilling by the Cordex Syndicate and Xxxxxx Mining Company from 1970 to 1996. This program led to the initial discovery of the A, and B pits, and led to successive discoveries of the C, CX, and Mag Pits. Within the Atna-Xxxxxx Mining Company Area of Interest there are 2,323 holes, totaling 1,063,734 feet, drilled by the Cordex Syndicate or Xxxxxx Mining Company.

4) A 206-hole program by Homestake Mining Company from 1997 to 2000.

5) A 3 hole program by Barrick during 2003.

6) Atna Resources completed 31 drill holes, a combination of reverse circulation rotary and diamond drilling from August 2004 through February 2005 totaling 29,740.5 feet of drilling

7) Miscellaneous exploratory holes in the area include a twelve hole program by Echo Bay Exploration, scattered holes by the USGS for research purposes, and numerous drill holes in and around the Granite Creek and Pacific Tungsten Mines.

Homestake Mining Company’s activities from 1997 to 2000 were the last extensive effort to explore for gold on the property, resulting in the identification of deeper high grade gold resources in the CX and Range Front Fault Zones which have been the focus of Atna’s exploration and resource definition work.

10.4 Trenching and Channel Sampling

No trench or channel sample programs are known.

10.5 Geophysics

Numerous geophysical surveys have been completed during the course of activity at Xxxxxx. These include regional and detailed surveys. Regional surveys include gravity and aeromagnetics. Detailed surveys run were mostly electromagnetic in nature and include IP, EM, MT, and CSAMT surveys. A brief summary includes:

1) Airborne EM and magnetics by the USGS at ¼ mile spacing throughout much of the Xxxxxxxx Trend.

2) Ground based magnetics covering the CX zone completed in 1970 by the Cordex Syndicate.

3) Regional gravity surveys, both public and private, compiled by Homestake Mining Company in 1997.

4) Ground-based magnetic survey at the north edge of the MAG pit completed in 1998 by Homestake Mining Company.

5) Several generations of AMT (EM, IP, CSAMT) completed by Xxxxxx Mining Company.

6) Several CSAMT lines completed by Homestake Mining Company in 1998 to 2000.

7) Several EM lines completed by Homestake Mining company in 2000

Technical reports and data sets are available for these surveys. No interpretive reports have been located. The geophysical surveys previously completed at the project have not been utilized by Atna as part of its exploration efforts to define mineralization associated with the CX and Range Front mineral zones.

10.6 Underground Drifting / Evaluation

A small exploration drift was put into the upper “B” zone by Cordex Syndicate in the early 1970s for bulk testing. Results are unknown, as no report has been found in the Xxxxxx property archives.

11.0 Drilling

11.1 Summary of Past and Present Programs

11.1.1 Drilling by Earlier Operators

A total of 2,325 holes have been drilled inside the Atna-PMC JV agreement area. Xxxxxx Mining Company, or Xxxxxx’x predecessors - Rayrock Mines and the Cordex Syndicate - drilled the majority of those holes (2,100), while Homestake Mining Company drilled 206 targeted holes. Holes drilled by other companies are scattered about on the property, including 4 by Barrick, 12 by Echo Bay at the north end of the property, and numerous holes drilled prior to 1970. The digital database, initially established by Homestake Mining Company, had been set up to track only HMC data from 1997 on. All data prior to 1997 was lumped into “OTHER” in the Drilled_By field, negating any method for tracking who drilled any particular hole. Table 11-1 Summary of Previous Drilling, provides the digital record of the previous known drilling activity on the property.

Table 11-1: Summary of Previous Drilling.

Drilled By	# of Holes	Total Footage	Average Depth
HOMESTAKE	206	236,255	1147
OTHER (PMC or Others)	2119	1,072,320	395

The vast majority of PMC holes were development holes in and around the existing pit areas. There are over 1200 drill holes within the A, B, C, CX, MAG, CXW, Xxxxx, and Bluebell pit areas. All but 9 are either conventional or reverse circulation rotary drill holes, and most (778) were drilled vertically. All of the 9 core holes drilled by Xxxxxx Mining are in the B, C, CX, and Mag pit areas, and served one or more of the following purposes; stratigraphy, metallurgy, or deep tests on mineralized structures. A nearly complete set of original drill logs is available in the archives of Xxxxxx Mining Company at the Xxxxxx project site.

Homestake Mining Company drilled 206 holes on the property with the vast majority (146) focused on the CX and Range Front Fault systems, accounting for 134,000 feet of the total HMC footage. Forty of these holes were drilled either entirely as HQ core, or as reverse circulation pre-collars with HQ core tails. The rest, including pre-collared portions of core holes, were drilled by reverse circulation methods. All original HMC drill logs are available onsite.

Xxxxxxx Gold Exploration drilled three holes in 2003 to test targets identified subsequent to Xxxxxxx’x acquisition of Homestake Mining. Drilling tested the deep (>3,000 feet) extensions of the CX fault zone near its projected intersection with the fault zone controlling mineralization in the Mag Pit.

11.1.2 Drilling by Atna Resources

Atna’s program followed up the drilling by both Xxxxxx Mining Company (PMC) and Homestake Mining Company (HMC) that identified the mineral zones beneath the CX pit and along the Range Front Fault. As part of its due diligence to define a mineral resource, Atna drilled 31 holes, totaling 29,740.5 feet of combined RC and core, to test previously drilled mineralization in the two primary targets. Atna’s program had four-objectives:

1.   To prove existence of the mineralized zones in terms of grade and thickness, particularly in regions where the only information available was reverse circulation drilling.

2.   To develop continuity to the mineral zones by drilling between known intercepts, particularly in areas where intercepts were greater than 400 feet apart.

3.   To establish size limits to known mineralized bodies, especially in areas where marginal and potential economic intercepts coexisted.

4.   To obtain rock quality data on hangingwall, foot wall, and mineralized zones to plan future underground exploration, reserve definition drilling platforms, and bulk sampling programs.

Of the 29,740.5 feet of drilling completed by Atna Resources Inc., 13,000 feet in 13 holes were drilled into the CX Fault zone and 16,740.5 feet in 18 holes were drilled into the Range Front Fault zone. (The prefixes APCX- and APRF- indicate the CX and Range Front targets, respectively) Table 11-2 Summary of Atna Resources Phase I Drilling, summarizes the drill holes by depth, RC footage, and core footage. Both the CX and Range Front faults are NE striking, SE dipping faults. Consequently, all of Atna’s drill holes were oriented to the WNW, around 300 degrees azimuth, and angled from 45 degrees to 75 degrees. Coring began between 100 to 200 feet above the fault zones and terminated usually from 50 to 100 feet after the footwall zone of the fault was encountered.

Atna’s area of immediate focus within the CX Fault zone and southern portion of the Range Front Fault zone contains numerous shallow drill holes (see Table 11-1), but only 370 drill holes from PMC and HMC actually xxxxxx the fault zones. The majority of holes drilled by PMC within Atna’s Phase I exploration area have either been mined out, or are too short to xxxxxx the fault structure.

11.2 Drilling methods

Atna’s program utilized both reverse circulation (RC) and diamond core drilling (DDH) methods. Reverse circulation drilling was used primarily as pre-collars for diamond core tails. This was done to minimize costs in the barren material above the mineralized fault zones. Diamond drilling was utilized to provide better confidence in sample quality than can be expected from RC drilling methods, as well as to provide for rock quality calculations and better geologic definition of the structurally controlled ore zones for engineering and modeling purposes. The distribution of reverse circulation versus core holes is discussed in Section 17.3 of this report.

Table 11-2: Summary of Atna Resources Phase I Drilling

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	0000	000	0000
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

11.2.1 Reverse Circulation Rotary Drilling

The RC drilling conducted by Atna was completed by XxXxxx Construction & Drilling of Winnemucca, Nevada using a Xxxxxxx 1500 truck mounted drill, supported by a water truck, pipe truck, and service truck, and staffed by a driller, helper and sampler.

RC drilling was used principally for pre-collaring core holes. The pre-collar portion consisted of drilling through non- or weakly-mineralized rock and stopping at a known depth above target mineralization. After drilling, the hole was cased using 4 5/8 inch threaded steel pipe to maintain the integrity of the hole until the core rig arrived to finish the hole to target depth.

Drilling started out using a 5 5/8 inch hammer bit. The hammer bit used a pounding action along with rotation to break the rock into pieces, and air pressure to lift the cuttings up out of the hole, and into the sample splitter. When the air system was no longer powerful enough to drive the hammer in the presence of high downhole water pressure, the hammer bit was replaced with a tri-cone bit. A tri-cone uses three rotating wheels with pointed carbide buttons to grind the rock into little pieces. Air pressure is used, as with the hammer bit, to lift the ground-up cuttings to the sampling apparatus.

11.2.2 Diamond Core Drilling

The diamond core program conducted by Atna was completed by two contractors, Xxxx Drilling of Xxxxx Fork, Idaho and Xxxxx Xxxxxxxxxxx of Chandler, Arizona.

Xxxx used an LF-1500 truck-mounted drill capable of depths up to 1500 feet with HQ rods, and 2000 feet with NQ. The rig was suitable for the shallower core tails, up to about 1400'. For deeper holes (up to 1800') it was necessary to resort to a Xxxxxxxx 3000 truck-mounted rig, capable of drilling to 3000’ with HQ rods. This rig was provided by Xxxxx Xxxxxxxxxxx. Both contractors staffed their rigs with a minimum of two (driller and helper), and added a water-truck driver for deeper holes.

Coring was completed using HQ triple and dual tube extraction methods. In triple-tube extraction, the rock is cut using a diamond-impregnated bit. The cut rock is fed through the center of the bit as drilling deepens and is stored inside a split tube contained in an outer core tube. After a run is completed, a wireline is dropped down the hole to pull the combined assembly. The triple tube is then extracted from the outer core tube by using hydraulic pressure to force the triple-tube from the outer tube. Core is recovered from this tube by opening the internal split tube and placing the core into the core boxes for transport, logging, and storage. This method ensures a greater degree of rock integrity, allowing for better lithologic, structural, and rock quality designation (RQD) descriptions of the rock. Standard dual- tube extraction of core follows the same procedure, except that the core retrieval is accomplished using just the outer tube of the above configuration, and is extracted by direct pumping of the cylinder of drill core itself, or by pounding the rock out of the tube. In highly fractured zones this often leads to degradation of intact core.

11.3 Logging

11.3.1 Reverse Circulation Rotary Chip Logging

Reverse circulation chips were collected by the drill sampler in 20-compartment plastic trays with recloseable lids, and brought to a logging trailer for examination under a standard binocular microscope. Each compartment contained a representative amount of drill cuttings from a 5-foot sample run. Gross lithologic types, alteration and other features were noted on the logs. A schematic graphic log was also drawn to aid the reader in interpretation of gross lithologic variations. Breaks in lithology and/or alteration were noted by drawing lines across the log sheet at the appropriate footage(s), with the depths of the breaks noted at the line. Sample numbers were plotted alongside the appropriate footages on the logs as an aid in comparing lithology and alteration to assays.

11.3.2 Diamond Drill Core Logging

At the logging facility, the core was laid out, left to right, on tables for examination. The core was oriented up-and-down (toward and away from the viewer), with the shallowest interval to the lower left and the deepest interval to the upper right. Logging tools included bristle brushes & spray bottles (to clean core), protractors, scribes, dilute hydrochloric acid, hand lens, binocular microscope, and logging forms.

The logging form includes a logical, visible place to record footage. Each length of core marked off by existing core blocks were marked on the log by drawing a horizontal line across the appropriate part of the log. The footage cut and recovered figure prominently in this record, with entries made for each "box" created by the horizontal lines marking footage. Intervals of no recovery were indicated by horizontal lines crossing the entire page, and a blanked-out zone of 'no information.' Visually, it is instantly obvious on the log what is missing and, perhaps, why. This is an appropriate place to confirm, add or incorporate the RQD data acquired earlier if it is not being maintained in a separate worksheet.

The logger pays particular attention to accurately documenting anomalies in footage shown on the blocks, and accounts for missing core by inserting blocks reflecting that, along with a cursory explanation on the block and a more substantial explanation on the log itself.

A graphic drawing of the lithology is present, to include major rock types using conventional or agreed-upon symbols, and the major structural features of contact relationships, dip and fractures, accurately plotted as to angle from the core axis. Other details of alteration and secondary mineral suites are added, as appropriate, and a comment area of the log includes a description of the rock that was cut.

12.0 Sampling Method and Approach

The objective of the sampling program is to collect a clean, uncontaminated representative sample that is correctly labeled when drilled and logged, and accurately tracked from the drill rig to the assay lab, and back to the mine site.

12.1 Sampling Methods

12.1.1 Reverse Circulation Rotary

During the drilling process, cuttings from the bit are sent up the drill pipe and initially into a cyclone for homogenization and mixing. From the cyclone, cuttings are fed into a rotary splitter that takes a representative split (usually a ¼ split), sending one split portion to the sample port, and the larger through the reject port. Cuttings are placed in 10” x 17” sample bags that are clearly marked using the drill hole number and a numeric sequence prepared beforehand using a spreadsheet. An example of the sampler’s logging sheet is included as Appendix 12-1, Sample Sequence). This sheet is used to track bag numbers and footages, standards, blanks, and duplicates. A small portion of sample is also kept for logging purposes and is placed in a chip tray compartment that is clearly marked as to footage and sample number.

12.1.2 Diamond Core Drilling

At the rig, core drillers are responsible for obtaining a complete and representative sample of the cored interval, generally in runs not to exceed 5 feet and in shorter increments in difficult conditions. Core is recovered from the barrel by using a wireline core tube, if possible outfitted with an inner 'triple-tube.'

12.2 Sample Quality - Recovery

Sample recovery for reverse circulation drilling is measured by weight of material captured. For a typical 6 inch diameter hole this will usually result in 8 to 10 lbs. of material on a ¼ split. Core recovery is measured by the ratio of length of material returned in the tube versus the total length drilled for the run, and expressed as a percent.

12.2.1 Reverse Circulation Rotary

Reverse circulation sample recovery was excellent, with full 5-10-lb. bags collected from every interval of every hole, with the exception of about 15 samples in the entire Atna set of approximately 6100 samples. The missing samples occurred in isolated zones of badly broken ground.

12.2.2 Core

Core sample recovery was also excellent, in excess of 99% of the 10,000 feet cored. There were fewer than 60 instances of core loss, each one averaging less than 2 feet and the vast majority of these losses were due to voids within the stratigraphy.

12.3 Sample Interval

12.3.1 Reverse Circulation Rotary

The normal truck-mounted reverse circulation drill in Nevada uses 20’ drill rods, and the sampler collects 1 sample every 5’. Such was the case with the Atna drill program. Samples were submitted for assay, as collected on the rig, in addition standards, blanks and duplicates were inserted into the sample sequence as described below in 12.5, 12.6, and 12.7.

12.3.2 Core

The typical truck-mounted core rig in the U.S. may be capable of using a core barrel up to 10’ in length, but clients will often require the core length to be limited to 5’ runs for reasons of sample integrity and to guarantee more onsite attention to the hole. The Atna program used a 5’ barrel. All of the core drilling completed by Xxxx - that is, the shallower holes - was recovered with a triple-tube assembly, which allows extraction of core from the inner tube with less disturbance of the core. A few of the deeper holes in the program, completed by Xxxxx Xxxxxxxxxxx, were cored with a standard two-tube assembly, because triple-tube equipment was not available.

After the core is logged, it is the geologist's responsibility to determine the appropriate sample intervals. As Xxxxxx is an underground exploration target, the geologist is careful to extract as much analytical information from the core samples as possible by adhering to rigid guidelines to better define boundaries between likely-mineralized and likely-barren samples. The original core blocks used by core drillers to mark the end of a cored run ordinarily serve as the primary sample boundary, subject to the rules below; where a conflict exists between the blocks and those rules, the rules prevail, and extra blocks are inserted by the geologist to compensate:

·   A sample must NEVER cross a lithology boundary.

·   A sample must not cross an obvious alteration boundary, including oxidation.

·   A sample must not exceed 7 feet in length, and only be that long if sure to be barren; 5 feet maximum is better.

·   Any core blocks that do not mark a sample boundary, for whatever reason (such as 'cave,', 'loss,' 'void,' etc.), must be labeled in black magic marker for photographic visibility.

Each block that marks a sample boundary is outline-highlighted in red magic marker, and these interval boundaries are entered on the sample sequence log alluded to earlier.

12.4 Sample Preparation, Quality Control Measures and Security

12.4.1 Sample Preparation and Quality Control Measures - RC Rotary Drilling

For each drill hole requiring reverse circulation drilling, the drill crew was provided with a set of bags prenumbered in sequential order (1 through XXX). The bags themselves carried only an exterior designation of drill-hole number and sample number. Since the sample numbering sequence includes blanks and standards inserted every 10^th sample, the driller's sampler cannot be expected to keep track of his sample collection based on the bag numbering. (Conventionally, RC exploration either relies on direct footage, or sample numbers in multiples of 5 feet). In order to prevent numeric confusion, yet permit Atna to submit "blind" sample numbers to the assay lab, several steps were taken.

First, the rig sampler was provided with chip trays accurately numbered with both true footage and the corresponding bag number. Second, he was provided with a deliberately incomplete set of bags; he was deprived of all the bags intended for standards and blanks. Third, since the ultimate completed depth of the hole is not known in advance, the bags for duplicates (to be collected every 100') were merely prelabeled with the letters 'A,' 'B,' 'C,' etc and flagged with a tear-off paper tag. The cuttings and tray chips themselves were collected as a continuous fraction of the return stream from the drill rig. The cuttings were diverted to a 10"x17" mesh bag, and the tray chips were diverted to a kitchen strainer. The filled chip trays were collected by an Atna geologist for logging under a binocular microscope and remain with Atna, while the sample bags are shipped to the analytical laboratory for preparation and assay.

Sample bags are allowed to dry/drain at the drill site and an Atna geologist visits the drill site and confirms the numbering and accuracy of the sample suite. Samples are then brought down to the shipment staging area, Atna personnel then relabeled the 'A,' 'B,' 'C,' etc. bags (representing the 100' duplicate samples) by assigning the correct sequential numbers for them; they form the tail-end of the sample list submitted to the lab and therefore are blind to the laboratory personnel. The samples are then loaded in 4' x 4' x 3' wooden crates for pickup by the laboratory.

12.4.2 Sample Preparation and Quality Control Measures - Core Drilling

Traditionally, core is forced out of core tubes by upending the tube, and tapping on it with a ball-peen hammer. Exceptionally, to preserve structural integrity, companies have requested/required that drillers attach a water line to the cap of the core tube and gradually "pump" the core out. This method is often ignored by the drill crew in the absence of close supervision. The triple tube core barrel eliminates this problem by allowing the interior core container to be opened lengthwise. Atna used a triple-tube for all of its shorter holes completed by Xxxx Drilling.

The core so obtained was carefully pumped out of the tube assemblies and laid out on a rack, intact. The drill crew recorded the Rock Quality Data (RQD) values on a worksheet and photographed it, using a digital camera to capture both the core and a whiteboard showing hole number and footage interval on display. For the deeper holes that did not receive triple-tube treatment and were therefore emptied in the traditional way, the geologist later recorded the most credible RQD values from core in the core box, and did not photograph it for RQD purposes, since it has already been 'broken up' for boxing.

The RQD measurement was established by measuring the total length of all pieces in a core run that exceed 2x the diameter of the core. In practice, 2x the core diameter is 0.3 feet. This number was summed, and written as a numerator in a fraction where the denominator was the total run drilled. On worksheets maintained by Xxxx, the raw values recorded are the length of the longest piece, and the sum of all qualifying pieces within the run Dividing the former by the latter yields the numeric RQD.

The drill crews placed the core in waxed cardboard core boxes, with tops and bottoms accurately labeled as to Company - Property- Hole ID - Box # -- From - To. The bottom of the core box was laid out longways from left to right, with the marked or labeled end to the left and the unlabeled end to the right. There are 5 rows or trays. The first portion of core was laid in the upper left-hand tray, and continuously laid in the tray from left to right, advancing "down" one row as each tray is completed. The bottom of the core terminated in the lower right corner. A wooden block was inserted at the end of each run, and in locations deemed important by the drillers to note adverse conditions, such as caving, voids, or mislatches (situations where the core tube failed to seat properly in the core barrel). The ending block for the run was marked with an ending footage on the thin edge, and two numbers on the larger surface:

C [cut] - m.n feet R [recovered] - m.n feet

The Cut number results from measured rod footage and the Recovered number stems from a taped measurement of core in situ.

If the drillers were not photographing the core, they marked the mechanical breaks they made to fit the core in trays with the letter 'M' on each side of the break. After boxing, the core was rubber-banded, a box at a time, and loaded into vehicles for careful transport to the logging area, where it was carefully unloaded and logged).

Each box of core, once logged (see 11.3.2, Diamond Drill Core Logging), was moved to the deck of the logging trailer for photography on a wooden stand with a digital camera, along with a legible placard indicating Hole # and From - To footage. After photography, the labeled end of the box top is marked on the upper right corner with a large red "P" in Magic Marker.

Since digital cameras were used for this photography, the quality of the photos was checked immediately before the core was disturbed for sampling. Both the core box photos taken by the geologic staff and the RQD photos taken by the drill crews were downloaded as quickly as practicable, checked for quality and clarity, renamed with meaningful file names, and printed out as 3x5 photos for archiving in 3-ring binders.

The geologists provided the sampler with working copies of two documents: the geologic core log and the basic sample sequence list, which contained the drill-hole number and a continuation of the numeric sequence carried forward from the precollar portion of the hole (Appendix 12-1, Sample Sequence). Although the sampler worked from the sample list itself, it was useful for them to be able to see why and how sample boundaries were picked. It also added a redundancy check on the geologist's accuracy.

A decision to saw versus split the core hydraulically depended on the condition of the rock and the geologist’s opinion of whether the core was barren or mineralized. Barren core was generally split with a hydraulic splitter, for the sake of speed. Mineralized silicified core was sawn; mineralized intact unsilicified core might have been hydraulically split, but this was not common (the vast majority of the mineralized core was sawn); mineralized broken core was invariably separated and divided equally.

The sampling operation avoided bias, wherever possible, by dividing the core in half perpendicular to the trace of the visible bedding. The portion to be saved remained in the core box, in its proper position, with core blocks in place, and the box was rubber-banded once again for safety. After the core was split, the samples were bagged and boxed in 4' x 4' x 3' wooden crates. Once logged, photographed and sampled, the core was palleted, covered with fitted tarps, and moved to industrial shelving on an outdoor cement pad for storage and reference.

12.4.3 Security - Reverse Circulation and Core Samples

Crated samples are delivered to the analytical laboratory in the numbered bags, along with a transmittal sheet stating whether the samples are “cuttings” or “core”, the range of sample numbers, and the total sample count. The lab has no knowledge of the spatial reference of the individual samples, beyond being able to figure out that sequential numbers from a drill hole represent top-to-bottom sampling. In the case of cuttings, they can also infer that the sample intervals are 5 feet long (standard in Nevada). In the case of core, it will be obvious from the volume that the maximum sample length would be 6 feet, but there would be no way of identifying any interval, and many such core samples will have a variety of lengths, ranging from 1 foot to 6 feet.

In addition, because of Atna’s insertion of blanks and standards in the sample stream, the lab cannot know with certainty exactly what footage a particular sample represents. Although forewarned that duplicates are present, the lab does not know where they occur in the group. By inspection of the submitted sample bag, the lab will be able to identify the blanks (red landscaping stone) and standards (pulp powder in Kraft envelopes), and will know that they occur in sample numbers divisible by 5, but will have no idea of the accepted value of each.

12.4.4 Sample Preparation

12.4.4.1 On-site

Sample bags that are intended as Standards and Blanks are labeled at the logging trailer, but these particular bags are removed from any numbered sequence bags provided to an RC driller at the rig to minimize errors. For standards, the code number and grade value is written on the sampler’s sheet at the correct sample ID value, and inserted into the appropriately marked sample bag.

Standards and Blanks, numbered as described above, are inserted in the crates at a rate of one standard and one blank for every 20 drill samples (see 12.5 and 12.6, below).

Each crate contains the raw samples, duplicates, standards and blanks intended for each hole. To the extent possible, all samples for one hole are aggregated together, and sample transmittal sheets are filled out in duplicate (one to the lab, one for file retention), with one job number assigned to each hole shipment.

Atna's copy of the transmittal sheet is stored in a three-ring binder in the logging trailer. Once assays have been received, a copy of the assay sheets will be stored with the drill logs and the original with the transmittal sheets. The transmittal sheets are indexed by job number.

Copies of the sample sequence list, the lithology log and assays are stored in three ring binders, indexed by hole number. Originals of all logs and assays are stored in file cabinets on a per-hole basis, also indexed by hole number. Atna personnel contact the lab to obtain a job number assignment for hole or partial hole shipment, and arranges for sample pickup by the lab's driver. In a number of cases, an Atna geologist returning to Reno on break may deliver a crate directly to the lab.

12.4.4.2 Laboratory Sample Preparation

BSi Inspectorate Analytical Laboratories, an ISO 9002-accredited facility (#37295), is the primary assay lab for Atna’s Xxxxxx Project analytical work. Sample preparation procedures employed by Inspectorate are as follows:

First, samples are thoroughly dried prior to crushing. Crushing consists of a two stage process. Initially samples are sent through a jaw crusher, and then through a roll mill to reduce better than 80% of the sample to -10 mesh. A 300 gram split is obtained from the coarse reject using a Xxxxx xxxxxx splitter. The split material is further reduced, with better than 90% of the split reduced to -150 mesh, using a ring and puck pulverizer.

After pulverization the sample is sent to the analytical portion of the lab where a 30 gram sample of the pulp is digested and analyzed for gold using standard fire assay methods. Samples are finished using AA, and, for any sample running over 3 g/t (0.1 oz/t), a gravimetric analysis is also completed.

12.5 Certified Standard Insertion

To increase the integrity of the sample handling process, from collection to shipment to assay, standards are inserted in the sample stream at a rate of one standard for every 20 drill samples. For each group of 20 drill samples, the 15^th sample is a standard: standards are therefore numbered 15, 35, 55, 75, etc. The reference standards were obtained from RockLabs in Elko, and consisted of a suite of both oxide and refractory certified powders. RockLabs also supplied statistics for the certified standards, shown in Table 12-1, RockLabs Reference Material.

Table 12-1: RockLabs Reference Material.

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

Certificates for the standards are provided as pdf files in Appendix 12-2 through 12-9.

12.5.1 Protocol

Standards are stored in plastic bins at the logging trailer. An appropriate powder from RockLabs is weighed in lots of 100g and placed in unlabelled, sealed Kraft envelopes. The envelopes are then placed in plastic bins individually labeled with the code number and gold content. The actual Kraft envelope containing the powdered standard is NOT labeled.

When a sample shipment is being prepared, standards with different gold concentrations are selected randomly from the available group in Table 12-1 and inserted into the sample stream. A listing of code numbers and corresponding values is added to each internally maintained sample list, archived, and available to all Atna personnel who deal with sample collection and shipping.

Standards were evaluated based on their variance from the accepted value, and by comparison to accepted standard deviations. If any assay from a standard exceeded 2 times the accepted standard deviation, the standard was flagged. If more than two of the same standard in the same batch failed, the batch was flagged. Comparisons to variances from values were checked continuously for the standards. Any batches where standards exceeded 10% variance were flagged and the assay lab was notified of the issues.

Standard assays are maintained in the database table, Appendix 12-10, CHECKS_STANDARDS

12.5.2 Summary of Results

Standard evaluation was broken into two sets based on the issues outlined in section 12.8.2. These issues are very apparent in the AA standards analyses, but gravimetric assay comparisons to the standards used are excellent (Figure 12-1). Out of 218 standards there were only 18 failures, or a rate of 1 in 12 (Table 12-2). Two of these failures were either misnumbered bags, or standards placed into the wrong bags (APCX-224 215, and APR-210 075), and 4 can be considered actual bad assays (APRF-227-035, APR-210-115, APRF-215-015, and APCX-214-075). All other assays have variances falling between 10 and 20%, with 5 standards assaying at just over 10%. An interesting observation for these standard results is that contained gold concentrations of the standards are between 3 and 8 ppm. All have exceeded the 2x accepted standard deviation flag, however, and have been considered failed.

Variances and standard deviation checks for the 110 AA finish standard results are poor. At least 70% of the samples failed the 2x standard deviation flag, and at least 40 exceed the 10% variance limit. Some of this can be accounted for. At least 6 standards are affected by out-of-sequence errors. These have been fixed. Others include two mislabeled standards (APC-206-035 and APRF-217-135), and six true busts.

There are 2 standards, OXK 18 and OXN 33, which routinely assayed with variances better than 10% (7 analyses) but missed the standard deviation flag. The remaining analytical problems involve the standards SF12 and OXE21. Accepted values are 0.891 ppm and 0.651 ppm, respectively. SF12 is a sulfide matrix while OXE21 is an oxide matrix. Standard deviations for these two particular standards are tight, at 3% each. Except for the 6 true busts all values generated for these two standards fall within 20% variance, and often between 10 and 15% of the accepted value (Figure 12-2).

The lab considered the values acceptable, according to its equipment standards. Since only two standards are affected by the phenomenon, matrix problems (improper digestions or fluxes caused by non-, or weakly-reactive, reagents) may be the issue. ALS Chemex has experienced issues with clay-rich matrices causing incomplete digestion (Xxxxxx Xxxxxx, personal communication, 2004).

Given the problems with standards SF12 and OXE21 it is hard to accept the values at lower levels. However these standards only represent assays from 500 ppb to 900 ppb, at the lower end of the scale of Atna’s current interests. Gravimetric analyses and AA analyses on standards over 3000 ppb are acceptable and lend confidence to the overall quality control program, particularly at the assay concentrations of interest to Atna.

Table 12-2: Failed Gravimetric standards

Job	Sample	AU_PPB	Au oz/ton	Accepted value	PPB Variance	Oz/ton for Standard	Oz/ton Variance
000-00-00	APCX-226 275	3609.00	0.12	3463	4.22%	0.10	16.83%
000-00-00	APRF-229 055	3438.00	0.12	3463	-0.72%	0.10	16.83%
000-00-00	APCX-224 335	3718.00	0.12	3463	7.36%	0.10	16.83%
000-00-00	APCX-216 215	6680.00	0.19	7378	-9.46%	0.22	-13.10%
000-00-00	APRF-227 035	20356.00	0.26	30490	-33.24%	0.89	-70.76%
000-00-00	APCX-224 215	18280.00	0.49	7378	147.76%	0.22	129.56%
104-27-01	APR-210 115	8190.00	0.29	7378	11.01%	0.22	33.83%
000-00-00	APCX-220 215	3900.00	0.11	3463	12.62%	0.10	10.89%
000-00-00	APRF-217 035	25821.00	0.77	30490	-15.31%	0.89	-13.64%
000-00-00	APCX-216 035	28550.00	0.69	30490	-6.36%	0.89	-22.86%
000-00-00	APRF-215 015	7000.00	0.21	8376	-16.43%	0.24	-14.86%
000-00-00	APRF-215 035	3690.00	0.11	3463	6.56%	0.10	10.89%
000-00-00	APCX-218 055	7765.00	0.24	7378	5.25%	0.22	10.60%
000-00-00	APCX-214 075	6596.00	0.21	8376	-21.25%	0.24	-14.86%
104-24-99	APR-211 115	3810.00	0.11	3463	10.02%	0.10	10.89%
104-24-60	APR-210 075	3541.00	0.10	7378	-52.01%	0.22	-52.60%
104-23-93	APR-208 155	7620.00	0.10	8376	-9.03%	0.24	-59.89%
000-00-00	APC-202 015	6720.00	0.19	7378	-8.92%	0.22	-10.78%

Figure 12-1. Standards Run With Gravimetric Finish.

(Red lines represent 10% variance)

Figure 12-2. Standards Run With AA Finish

(Red lines represent 10% variance).

12.6 Blank Sample Insertion

Blank samples are used for the purpose of checking the primary and secondary crushing process at the analytical lab, particularly the adequacy of equipment cleaning between samples. Blanks are considered to have failed if they exceed 2 times the detection limit of the analytical device. Since Atna has not used pure blank material, blanks are considered to have failed if they exceed 20% of the accepted values as outlined above. Assay are to be re-run if failure rates exceed 1 in 5 samples.

12.6.1 Protocol

Commercial decorative stone purchased commercially in 50-pound bags was used as the blank material. The material was stored, in the original packaging, in a bin outside the core-logging trailer. One to two 18 oz scoops of material per sample was considered sufficient (about 1.2 kg per sample). The material was placed in sample bags marked with the appropriate hole number and sample number for insertion into the sample stream, then stored with the actual drill cuttings prior to shipment. Blanks are inserted into the sample stream at a rate of one blank for every 20 drill samples. For each group of 20 drill samples, the fifth sample is a blank: blanks are therefore numbered 05, 25, 45, 65, etc'. Blank samples are inserted into the sample sequence on site and prepped by the lab with the actual drill samples.

A series of initial assays were completed on the decorative stone to provide a base line for analytical values. These results indicate the material, although not completely devoid of gold, is sufficiently barren to be able to discern if contamination occurs because of improper cleaning techniques at the lab. Two separate pulps were created for the gold assays. Table 12-3 lists the analytical results.

Acceptable values for the material used to evaluate lab sample preparation methods for drill assays are from the clipped column, where the highest value and lowest value were thrown out. Ranges of acceptable values are from less than detection to 26 ppb (1 standard deviation from the mean). Samples exceeding the upper end of the range are considered to have failed.

12.6.2 Summary of Results

A total of 358 blank samples from 63 separate assay jobs were compared for failure rates. Of the 358 samples only 21 failed, yielding a failure rate of 1 in 18 for the entire program. Of the 63 assay jobs analyzed, only three jobs contained samples exceeding the 1 in 5 failure rate for inserted coarse blanks.

Seven blank samples from the failed jobs were reviewed. Of those seven, 2 samples represented poor cleaning, and the remaining 5 are unknown failures. The two samples representing poor cleaning were preceded by high grade results (>0.100 oz/ton) and came from the same job. Two other blanks from this job were also preceded by high grade samples and passed the failure rate screen. These two samples represent an isolated incident.

The 5 failures in question were preceded in sequence by normal drill assay samples with gold values less than the blank samples had, and may indicate contamination from another step in the assay procedure. However, all other samples surrounding the blanks assayed with consistent low grade values and showed no signs of contamination spikes. It is entirely probable that the blank material used has greater variance in gold concentration than the initial round of assays on the material would suggest.

Results for the blank insertion program indicate proper cleaning procedures were followed at the lab, and sample preparation errors have not influenced assay results.

Table 12-3: Decorative Stone (Blank Sample) Analysis

Decorative stone Initial analysis - September 9, 2004
BSi Inspectorate Final Report - Job No: 000-00-00
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

12.7 Duplicate Samples - Reverse Circulation Rotary

Duplicate samples were taken in the field at the drill rig every 100 feet and were used to evaluate sampling protocols at the rig to ensure adequate representation of material was being obtained. Duplicate samples should fall within 30% of the original assay with a failure rate not exceeding 1 in 10 samples. If the failure rate exceeds 1 in 10 samples then collection methods at the rig need to be modified.

12.7.1 Protocol

Duplicate samples were taken every 100 feet of depth at the drill rig, with the first duplicate sample starting at 95-100 foot depth. Samples were obtained by placing a 50-50 splitter on the sample port of the rotary splitter when drilling wet. Dry drilling required the use of a Xxxxx splitter where the samples, both assay and duplicate, were obtained from a 50-50 split of the sample material from the cyclone.

Duplicate samples were inserted at the end of the sample sequence. In order to maintain a continuous sequence of sample numbers a second set of sample bags were labeled with hole number and “Duplicate A”, “Duplicate B”, etc.. Attached to the bag was a removable piece of paper with hole number and footage for the duplicate sample. This labeling system was necessary because the last sample number at the end of the RC portion of drilling was not known until finished. After RC drilling was completed, new bags for the duplicate samples were labeled that continued the proper numeric sample sequence from the end of the hole. Duplicate sample material was then placed into the properly numbered sample bag for shipment to the lab. Records were kept on the sampler’s record sheet noting the footages for each duplicate sample and a cross reference from the primary bag label (Duplicate A, etc.) to the actual sample number.

12.7.2 Summary of Results

Figure 12-3 shows is an x-y pair graph showing a 30% warning line from a baseline reference. Samples falling on the high line exceed the 30% acceptance limit and therefore have failed the protocol screen. Of the 171 samples only 5 fall above the line. This yields a failure rate of 1 in 20.

Two of the five samples were determined to have had the bags switched, either at the rig by the sampler, or during the relabeling of the bags for insertion in the sample sequence. It has not been determined which has occurred. The other three samples are due to poor quality control at the rig, with a poor split as the likely cause.

Results from the duplicate program indicate sampling procedures at the rig are sufficient to indicate proper representation of the rock sample is being sent for assay.

Figure 12-3. Comparison of Duplicate Samples Taken at the Drill vs. Xxxxx Xxxxxx.

12.8 Check Assays of Mineralized Samples

Check samples were sent to ALS Chemex labs for verification of assays run by the primary lab, BSi Inspectorate. A total of 214 samples were sent for check analyses. Check assays should fall to within 10% of the original value if the original pulp was used for the re-assay, and within 30% if the check assay came from a coarse reject duplicate. Failure rates for jobs occur where less than 90% of the samples do not fall within the 10% error rate.

12.8.1 Protocol

The check assay program consisted of identifying samples with initial results greater than 0.10 opt gold, and included samples with values less than 0.10 opt gold if they were part of a mineralized zone with an overall intercept at better than 0.10 opt gold. After identification, a job order was placed with the original lab (BSi Inspectorate) to have the original pulps pulled and sent over to ALS Chemex for re-analysis. All check samples were analyzed by ALS Chemex using a 30 gram split of the original pulp and run using standard fire assay techniques with a gravimetric finish.

12.8.2 Summary of Results

A total of 318 samples were submitted to ALS Chemex as checks on BSi analytical quality. At first glance the number of samples exceeding acceptable error limits is over the allotment given for failure rates (Figure 12-4).

Figure 12-4 is a plot of sample variance, sorted by grade with higher grades (values in ppb) to the right. The red line is the 3000 ppb demarcation point, and the green line is the 150 ppb demarcation point. Samples below 150 ppb Au show extreme variations, which is understandable at such low concentrations. Samples between 1000 and 3000 ppb gold show more consistent variation, with the majority falling within a 20% error limit. In samples above 3 ppm the variations become much tighter, with the majority falling into the 10% error range. The reasons for the variance are the graph essentially compares AA finishes from BSi Inspectorate with gravimetric finishes from ALS Chemex. To truly check the validity of the check samples one must separate the data into two sets, and eliminate all samples below 150 ppb Au due to the inherent high variations generated at such low concentrations of gold.

Figure 12-4. Sample Variance of All Check Samples

Eliminating samples with values below 150 ppb Au leaves 254 samples. Of these 254 samples 110 represent assays between 1000 and 3000 ppb Au comparing gravimetric to AA finishes, and 144 above 3000 ppb Au comparing gravimetric to gravimetric finishes. Figure 12-5 shows the checks versus original assay pairs and a 10% error level (redline) relative to expected values for the 254 samples above 150 ppb Au (this graph is also a mix of AA-gravimetric comparisons). Table 12-4 summarizes the results of the two grade intervals.

Table 12-4: Statistics of Lab Variance in AA vs. Gravimetric Finish

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

The graph shows 56 samples to occur at or above the 10% error level but only 12 samples well above the 10% error level. The rest of the samples appear to hug the line, falling between 10 and 20% error levels. Since the graph really is representing AA versus Gravimetric analyses the errors between 10 and 20% can be explained by measurement errors associated with AA analyses above 8000 ppb, and measurement errors associated with gravimetric analyses below about 5000 ppb (errors for both methods become larger as the upper and lower thresholds of the equipment are approached and exceeded). This would tend to exacerbate the errors in variation seen between the two methods.

Figure 12-5. X-Y Pair Plot of Check vs. Original Assay, By Grade

To fully estimate the percentage errors associated with the two methods they must be compared separately. However, for samples between 1000 and 3000 ppb, the comparisons are still between AA and gravimetric finishes due to the analytical methods applied at ALS Chemex. Because these samples may fall below cutoff grade it is not as critical to compare these explicitly. Figure 12-6 shows the results of the samples between 1000 and 3000 ppb Au.

The red line on the graph represents the 1000 ppb break, with samples to the right of the line assaying above 1000 ppb. Variations left of the red line consistently fall within the 20% error level, but a significant number falling outside the 20% error level also. Variation errors become much less as higher grades are encountered, and by 2000 ppb (green line) very few analyses have greater than 20% variance. As 3000 ppb is approached the variances become far less, and begin to fall within the accepted 10% variance level.

In contrast, assay variations on a gravimetric to gravimetric comparison are typically within the less than 10% error limits (Figure 12-7), even though a significant number also fall between 10 and 20% variance. The larger proportion of these samples assay at less than 8000 ppb, toward the lower end of gravimetric finish ranges and may be accounted for by errors associated with the gravimetric units at both labs.

Comparing the gravimetric results from ALS Chemex to the AA results from BSi Inspectorate for these samples one finds that sample variance between the two assays fall within the 10% accepted limit. This suggests BSi Inspectorate gravimetric machinery may not be well calibrated at lower gold concentrations. Other considerations such as differences in reagents and fluxes used, and experience of the fire assay personnel may also lead to unidentifiable errors between the two labs.

Figure 12-6. Variance of AA-Finish samples < 3000 ppb

Figure 12-7 Gravimetric Finish Comparisons - -Check Assays to Original Assays

Although there were many more samples outside the accepted 10% variance limit failure rate, only 14 serious busts were found that required attention (Table 12-5). These busts were from 3 check assay jobs, with only one job (RE04074554) considered as a bad assay batch, but from the check lab (ALS Chemex) rather than the primary, BSi Inspectorate. These 4 samples were re-assayed, with results bringing the samples into conformity with acceptable standards.

Table 12-5: Failure Rate for 3 Check Assay Jobs.

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

Of the remaining ten assays, 4 samples assay below 6000 ppb and can be attributed to machine error, the other 5 have values greater than 0.5 oz/ton. These higher-grade assays have variances in gravimetric analyses at less than 20%, usually between 12 and 15%. Discussions with the laboratories suggest differences in equipment, reagents, and methods may account for the variations. Variations at this concentration level, in so few samples, can be accepted because of the consistency in error between the two labs (the errors are always between 12 and 15%).

Due to the inconsistency in assay method applied it is difficult to assess the accuracy of the laboratories’ analytical procedures for the entire batch of samples. The simple fact that variance errors are consistently less than 20% for most comparisons indicates variation due to analytical methods (AA vs. gravimetric), rather than procedural errors, can account for the differences. The errors in gravimetric comparisons, the only truly correlative analytical methods that are available, suggest some tightening of procedures can be achieved at both labs (busts at Chemex and low gravimetric analyses at less than 8 ppm at Inspectorate). Only 14 assays, due to a variety of reasons, from a set of 254 were found to be of truly poor quality. These represent less than 10% of the number of assays used in the statistical studies, and indicate acceptable levels of analytical quality for the primary assay values.

12.9 Laboratory Quality Assurance and Control

Internal QA/QC samples were provided to Atna Resources Inc. by the assay lab. Internal lab protocol consisted of running 1 in 20 to 1 in 40 samples (depending on job size) as internal duplicates to check assay adequacy. Results were posted on the final certificate under the heading “QA/QC Au ppb”. The duplicate sample was taken from the coarse reject with the preparation of an entirely new pulp, and run as blind samples within the job. Evaluation of the results looked at the number of samples falling outside an accepted 10% range. If less than 90% of the samples are within the 10% range, then affected assays falling outside the 10% range should be re-run.

Evaluation of the laboratory duplicates compared the original assay to the blind duplicate. A total of 460 duplicates were inserted by the lab throughout the program. To make meaningful comparisons, though, only those samples greater than 30 times detection (above 150 ppb) were used, as variances at lower levels are extreme and meaningless. Given these parameters, only 177 samples were used to evaluate lab internal QA/QC. Of the 177 samples, 10 (5.6%) fell outside the accepted 10% of the original assay (94% of the samples were within the allowable 10% error range).

Figure 12-8 represents an x-y scatter plot with red lines representing a 10 percent error. Two samples with the largest variances were found to contain data entry errors. These have been corrected in the database. Other samples exceeding the error limits are unexplained, but given only 5.6% of the samples were bad, the internal lab duplicates validate assay adequacy and accuracy.

Figure 12-8. Laboratory Blind Duplicates

13.0 Data Verification

13.1 Summary

Verification of previous exploration drilling assay results is required as part of prudent due-diligence studies for the acquisition and ongoing exploration of the Xxxxxx Mine property. The process of data verification follows several threads. One thread is retroactive correction and maintenance of the large Xxxxxx Mining Company database. A second important thread consists of re-analyzing existing drill sample pulps from mineralized intercepts within the CX and Range Front target areas. Pulps were taken from the onsite pulp library maintained by Xxxxxx Mining Co., which includes all drilling done on the property while PMC was operator.

13.2 Database of Previous Drilling

Barrick supplied digital data on CD for the Xxxxxx Mine. Contained on the CD were two Microsoft Access databases containing surface sample locations and geochemistry, and, most importantly, drill-hole location, assay, geology, and survey data. This data set was also duplicated in a Vulcan data set, and as ASCII text files used to import/export in Vulcan. After Xxxx established a working office at the mine site, additional data was found, and it was learned that Homestake personnel had put together the entire data set digitally.

HMC essentially compiled the Xxxxxx Mine assay database under two separate efforts. One was the entry of older Xxxxxx Mining company data, probably from a combination of hand entry and extraction from the MedSystem database that was used at the mine. The other was the acquisition of new data from ongoing HMC exploration efforts. Data were initially compiled in Excel spreadsheets before final storage in an Access database. Homestake’s Access database contained tables for the following information:

1) Collar information

2) Downhole survey information

3) Geologic information

4) Assay information

5) Geochemistry

6) Grade thickness

While the database is useable in the structure it was created in, Atna had additional data to track, and decided to create a new data structure appropriate for its own work, and then import the Homestake data structure into the new Atna structure.

Data acquisition issues forced the datasets into 2 separate Microsoft Access databases, one for Atna Resources data and the other for historic data. Data management was structured in this manner to allow ongoing cleaning of historic data without interfering with introduction of newly acquired data. These data sets are being maintained as separate entities until they can be integrated with confidence. Both databases have the same table structures, field names, and table names, and should be easy to merge.

13.2.1 Tables

In both databases, the main data sets are contained in the following tables:

DH_Collar - all collar information

DH_Assay - all gold assay information

DH_Geochem - all drill-hole trace element geochemistry

DH_RQD - RQD data

DH_Survey - down hole survey data

DH_Geology - lithology and formation information

DH_Alteration - alteration data

For the Atna Resources database the following tables also apply:

TRANSMITTALS - lab job certificate numbers (key field for assays and a tracking device)

DH_SAMPLERS_RECORD_SHEET - cross reference list of drill-hole footages, sample numbers, standards, and blanks. Sample Type is also referenced so as to quickly discriminate between core from RC.

CHECKS_STANDARDS - all check and standard assays for QA/QC

Although these tables exist in the historic data set there is no information contained in them.

13.2.2 Data Corrections

Prior to entering data into the newly created data structure, any errors were identified and resolved. The basis for fixing identified errors was the well-maintained filing system containing most, if not all, drill logs, downhole survey data, and HMC assay data. While Xxxxxx Mining Company was operator, assaying was done in the mine’s own labs, and Atna Resources has yet to locate original drill-hole assay sheets from that era. Atna's only source for these gold assays are values hand-entered onto lithology logs. Consequently, it has been difficult to check the validity of some historic data, except as part of Atna’s reassay program completed on existing pulps from these earlier drilling programs.

The verification process involved the following steps:

1)   Comparing collar data to survey, assay, and geology data for incorrect depths, Azimuths, and dip values.

2)   Check survey, assay, and geology data for missing and/or overlapping intervals

3)   Randomly check drill-hole geologic, survey, and assay data against logs and hardcopy data.

4)   Plot sections and check for obvious errors in hole trace orientations and potential missing assay data (no data present).

The following errors were detected and resolved:

Collar Table - Approximately 100 known errors involving incorrect depths, azimuths, and dip values have been corrected

Geology Table - Approximately 70 known errors correcting length overruns (survey depth greater than TD), overlapping intervals, and missing intervals.

Accuracy of geologic data is subjective and based on the logger for lithology and formation call. Typographical errors in the database are not known, but probably occur. Other errors that have not been fixed include multiple entries (i.e. Flt, FT, Fault) for the same data type and duplicate entries for the same interval. Actual drill logs will be Atna’s definitive source for interpreting geology.

The geology table has not actually been loaded into the new data structure, but rather resides in its original form as a separate table within the database.

Survey Table - Approximately 110 errors including length overruns, missing intervals, out of sequence intervals, and bad azimuth or dip values.

Geochem Table - All data has missing intervals. Data has not been fixed yet, and will be done on loading when data is needed. Homestake did not run all intervals from the drill holes sampled in their geochemistry program, leaving large gaps in the from and to sequences within the database.

Assay Table (DH_Assay)

This is the most important table in the database for generating resource estimates and must be clean. The original structure of the table contained the following fields.

Table 13-1: Field Definitions in (HMC) Xxxxxx Database

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 Xxxxxx Fire Assay in oz/ton
AUCN	HMC cyanide assays and Xxxxxx XX 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

This structure was used primarily for pre-Atna resource

modeling efforts in Vulcan software.

Many of the problems and errors encountered in the assay database came about as a result of the pulp re-assay program. That program, described below, caught several errors, ranging from simple clerical entry errors to the practice of entering only single Fire Assay values for the Xxxxxx Mining Co. data, regardless of check assays and any additional assay reporting. Below is a summary of errors caught while validating the data.

Incorrect assays: - Several holes (for example, RH-123) have assay values entered showing grade, where the only source of information indicates either no grade, or assay values in the wrong interval.

Missing intervals: - Approximately 25 holes with intervals actually missing.

Intervals beyond hole depths: - Approximately 60 holes with intervals beyond hole depths, including many intervals containing anomalous and low grade gold values.

Overlapping intervals: - 35 holes with overlapping intervals.

Typos: -Several assays have typos caught by the pulp re-assay program. Many probably still exist. In one instance the typo changed a 10 ppm assay to 1 ppm

Other errors in the assay database that are less egregious include:

"Less than" detection: - Although not significant for ATNA's purposes, since the company is focused on high grade values, it leads to a general sense of poor quality control and checks on the original database. HMC converted all "less than" detection values to "+½" value. This was done for statistical purposes during resource estimation. Converting less than detection values is not an issue generally, but it has been inconsistently applied. All “less than” detection values have been identified and converted back to reflect original results on certificates or drill logs.

Missing assays: - Homestake Mining Company entered "-1" for any assay that was missing. This included -1 for any analyses that were not run, and for samples with missing intervals due to poor recovery, voids, or fractures. Since Atna is pursuing a structural target, it is important to make the distinction. In the current database a value of -8 or -9 has been entered to reflect the distinction. "-8" values indicate no assay was run, even though a sample was taken, or if it is not known why an assay is not present. A "-9" indicates no assay due to insufficient sample, or missing sample, attributable to geologic reasons. Numerous holes have been fixed but many others have not. These discoveries continue as holes are plotted on section.

13.2.3 General description of pre Atna assay results and procedures.

Because no assay certificates exist for Xxxxxx Mining Company drilling, it cannot be said with certainty where actual assays were completed, nor whether the assays were accurately entered on the log sheets. It has been generally accepted the assays were completed at the mine site.

Xxxxxx Mining Company's standard practice was to run assays using AA methods. For all assays this was generally done on a cyanide xxxxx, as interest lay only in identifying leachable ore. After a certain date (unknown) Xxxxxx Mining began to run fire assays with an AA finish on all assays over 0.01 opt, and selectively ran check assays at third-party laboratories for numerous high-grade zones. All the assay information was hand written on the lithologic log.

Xxxxxx Mining Co detection limits were <0.003 and <0.001 oz/ton, with the distinction between detection limits being the age of the assay. These values were also converted to “+1/2” detection limit values. In addition, 0.0005 and 0.0006 oz/ton were also found in the database for Xxxxxx Mining Company Holes. On the original logs, these assays were reported as “Nil”, <0.003, or as "<0.001 oz/ton." For Xxxxxx Mining company holes, all “less than” detection values have been converted to -0.001 or -0.003 for oz/ton in the database.

Validity for using Xxxxxx Mining Company data derives from the pulp re-assay program.

Homestake Mining Company assayed all drill holes at ALS Chemex in Reno, NV. ALS Chemex provided sample pickup and preparation services. These procedures are consistent with standard practices in the industry and include:

1) A primary crush to less than -80 mesh

2) Split of 300 grams of material for pulverization to -180 mesh

3) A 30 gram split for digestion and assay

All assays were completed using the AUAA23 Fire Assay method. Initial assay results were finished using an AA machine with results reported in ppb values. If results came back less than 10,000 ppb gold these were reported on the certificate using the actual ppb value. Samples that assayed greater than 10,000 ppb on the initial AA finish, were re-assayed using an assay-grade gravimetric finish. These gravimetric assays were reported on the same certificate using ounce per ton values in a separate column. The value “>10000” was placed in the ppb column for any gravimetric analyses.

Detection limits for gold analyses on ALS Chemex assay certificates are -5 ppb and -0.0005 for oz/ton. Most HMC holes reporting such values were converted, for statistical purposes, to 2.5 ppb and .0003 opt. Several HMC holes also had actual ppb results of 34.3, 17.8, 18.9, and 56.0 reported as 2.5 ppb values in the database. For HMC holes all "less than" detection values have been converted back to -5 and -0.0005, respectively, for ppb and oz/ton in the database.

Homestake also ran, as standard practice, a cyanide shake xxxxx assay on any result greater than 0.01 oz/ton (334 ppb). These were reported on the certificate under a column labeled Au CN OPT. Detection limits for the cyanide assays were less than 0.001 oz/ton.

Pulps from HMC’s exploration drilling were also subject to re-assay.

13.3 Pulp Selection

Atna Resources re-analyzed 652 pulps from drill-hole samples as part of a due diligence effort to confirm the presence of strong gold mineralization at depth and along strike from open pits and other hypothesized structures at the Xxxxxx Mine. Selection of pulps was based on grade, and whether or not the mineralized intercept was contained within either the Range Front or CX fault zone. Grade ranges for samples to pull were those intervals above 0.10 oz/ton, unless the sample was internal to a mineralized intercept. The pulps came from the mine’s pulp library stored on site.

Intervals pulled are listed in Appendix 13-1, Pulp Re-Assay List. Care was taken to maintain the order and status of the library. All pulps pulled for re-assay were replaced in their original boxes, in order, and the boxes restored to their original locations on the mine site. After pulling and cataloguing the pulps, Atna staff delivered the pulps to the BSi Inspectorate laboratory facility in Reno for re-bagging and insertion of blind standards and blanks. ALS Chemex Laboratories completed the analytical work.

13.4 Re-Assay Methods

13.4.1 Processing by BSi Inspectorate

BSi Inspectorate re-bagged the pulps using the following process:

1)   In order to maintain a uniform sample sequence with the insertion of blind standards and blanks original pulps must be re-bagged and re-numbered.

2)   Existing pulps were rolled and a 100 gram split taken for the new assay.

3)Pulps were re-bagged in a new envelope and given a new sample number. A cross-reference list was maintained to index the original pulp sample numbers to the new split pulp sample numbers. The original list was given to BSi Inspectorate as an Excel spreadsheet containing existing pulp sample numbers and locations for the insertion of blind standards and blanks. BSi Inspectorate then filled in an index field on the list with the new sample number sequence. The list was delivered to Atna Resources, both as hard copy and digital, once pulp preparation was completed.

4)BSi Inspectorate inserted certified standards from CDN Resource Laboratories Ltd. into the sample sequence at a rate of 1 standard for every 20 samples. Standards were given sample numbers that maintained a uniform numbering sequence for the entire job (see step 1). These numbers were listed on the cross-reference sheet with the following additional information: accepted assay value for the standard, matrix type, and Standard Reference ID #. Statistics for the certified standards were available from CDN Resource Laboratories’ website. Table 13-2, CDN Resource Laboratories Standards, lists the standards.

5)BSi Inspectorate also prepared and inserted into the sample sequence silica sand blanks at a rate of 1 for every 20 samples. Standards and Blanks were not inserted adjacent to each other. Again, sample numbers for blanks maintained a uniform numbering sequence for the entire job, and were identified in the BSi Inspectorate cross reference list as “silica sand blank” along with a corresponding sample number.

6)Samples were delivered to ALS Chemex in four separate batches for analysis.

13.4.2 Processing by ALS Chemex

Samples were analyzed by ALS Chemex with these conditions:

1)   New fusion crucibles.

2)   Gold assays determined by 1 assay-ton, Fire assay with an AA finish (Au-AA25), or via a gravimetric finish for samples greater than 100 ppm.

3)   Results reported directly to Atna Resources.

Atna compared and contrasted the reproducibility of the gold values obtained in the re-analysis project with those originally reported to Xxxxxx Mining Company. A statistical analysis of the analytical data, certified gold standards, and silica blanks was completed to assist in gauging the reliability of the data.

Table 13-2: CDN Resource Laboratory Standards

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

13.5 Re-Assay Results

New assay values for the pulp program came back from ALS Chemex in 5 separate jobs. ALS Chemex split the second pulp batch (491 samples) from Inspectorate into two separate jobs (270 and 221 samples) because the number of samples exceeded Chemex’s internal QC job size allotment of 300 samples. Results from those assays were compiled and statistically analyzed by comparing the new assay values to the original values.

Initial results of that analysis showed a significant decline in gold concentrations for numerous samples when compared to original assay results, many of which were also done by ALS Chemex Laboratories. The lower gold concentrations occurred at all grade levels, but happened to be particularly acute in samples assaying above 8 g/t Au. Additional statistical evaluations also showed 20% of the analyses fell outside of an accepted 10% variance from original gold results, as well as uniformly lower results and large variances in values of standards when compared to accepted values.

Several samples were pulled from each assay job for additional testing. Statistical analyses of results from the testing, with some exceptions, fell within accepted 10% variances from original assay results (Figure 13-1). This is particularly true of samples above 10 ppm, which, in the initial re-assay tests fell below the -10% variance level, and of all standards re-analyzed during the testing (Table 13-3). Re-analysis of these samples led to the conclusion that bad assays of the newly-submitted pulps were to blame, in particular improper fluxing of the sample. Two jobs had to be re-analyzed based on the tests.

Figure 13-1. Test Result Variances

(Red lines are 10% acceptable limits).

Table 13-3: Analyses of Standards Used in Testing.

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

* Note sample 27 was initially misreported as to original value (1775 ppb). The value should be 780 ppb.

High variations in results between 10 and 25 ppm from several samples in Figure 13-1 are from samples run by Xxxxxx Mining Company. These variations are at odds with accuracy comparisons with the standards analyzed in the same batch. Checks of the original logs traced discrepancies in the data to various problems, including database entry errors, variances in original assays, or poor quality control on older Xxxxxx Mining Co. jobs. Use of different pulps for fire assay analyses and check samples is also a part of the problem, as indicated by several analyses (e.g. RHA 456 460-465) where the Fire assay value (on a different pulp) was much less than the original AA assay, and Xxxx’s re-assays reflect the AA result. When comparing re-assay results with Xxxxxx Mining’s original data there is at least one original assay value that does correlate well with the re-assay results, indicating the pulp in the library came from this assay rather than the original AA assay. In any event, these variances were tracked down and an acceptable reason and solution for the variance was then determined.

After resolution of variance issues, the results were incorporated into the data set and re-analyzed statistically. The analyses fell within acceptable limits on the second assay set and have been incorporated into the database as the actual values for the re-assay effort.

13.6 Discussion

Atna’s re-assay results substantiate the accuracy of the original HMC and PMC assay results, showing the original numbers to be valid. Furthermore, the re-assay results also lend a degree of confidence to the Xxxxxx Mining Company assay data as a whole, indicating that the values obtained by Xxxxxx Mining’s original assays and the values written on the logs and entered into the database, can be used even for those holes not re-assayed under this program.

After verification and correction of database records, Xxxx was able to enter the data set into a newly structured Access database. The assay table was structured to reflect the true nature of the assay data, with fields for AA,, gravimetric, Xxxxxx fire assay data, and Homestake ppb data. A field was also added to store the values to be used in resource modeling and estimation.

Since there are several types and generations of assay data, even for a particular sample, the concern with the data set was, Which gold assay value to use for resource estimation. Standard geostatistical practice is to use the original assay so as to not alter the inherent variability within the deposit (Xxxxx Xxxxx, personal communication, 2005). An order of precedence must be used to ensure that the highest quality assay method, and thus the highest quality value, is used from the several generations of available data. The restructured Atna database uses this order of precedence:

1) Atna re-assay data for HMC and PMC holes

2) Fire assay gravimetric data for HMC and Fire Assay AA data for PMC holes.

3) Fire assay ppb values for HMC holes converted to OPT using - oz/t = Au ppb/34285.7

4) AA assays for PMC holes.

5) Cyanide assays for all others (purely theoretical, as some higher-quality measurement exists for every sample).

This precedence was applied and the data stored in the Field AuFA_Calc, and are the values to be used for all modeling and resource estimation purposes.

14.0 Database Audit

14.1 Analytical Data

A separate audit of the assay database followed the verification and cleaning process described in section 13 above. This audit was of assay data from an MS Access database of all drilling completed on the Xxxxxx Mine Property to the end of exploration activities in 2001, and consisted of a mix of data from Xxxxxx Mining Company and Homestake Mining Company drilling. Drill data of material interest to Atna are considered to be those holes intersecting the Range Front or CX fault zones in areas where identified gold mineralization has significant potential to be developed into economic reserves. A total of 370 holes from the Xxxxxx Mine Assay database, identified through cross-section work, are being used for modeling and resource estimates for these two zones.

14.1.1 Validation Process

Seventy-four drill holes, representing 20% of the 370 holes of interest to Atna, were picked randomly using Microsoft Excel:

1. Each of the 370 holes was given a unique numeric index number in the Access database and then exported to Excel.

2. Using the Data Analysis toolpack add-in for Excel, a random sequence of numbers was generated using the unique index number as a reference.

3. The selected index numbers generated by Excel were imported back into Access and cross-referenced to drill-hole ID.

4. An assay set was extracted from the database using the cross-referenced drill-hole number list. This list was further winnowed for checking purposes by examining only assays greater than or equal to 0.08 oz/ton.

5. Assay values from the database were compared to assay certificates for HMC data, and against values presented on drill-hole logs for Xxxxxx Mining Company data. (Actual assay certificates have yet to be found for Xxxxxx Mining Company work.)

This procedure generated a total of 7,479 individual assays for the 74 drill holes. Extracting values greater than or equal to 0.08 oz/ton yielded a total of 216 assays to be checked for errors. Appendix 14-1,: Error Analysis of Randomly Selected Xxxxxx Database Assays, lists the hole numbers and assay intervals used in the check process, and the errors found.

Of the 216 assays checked there were 16 errors, representing 7% of the total number checked. Only 7 of the 16 would materially affect the database (3%). Of the 16 errors, four represent check values entered that were inconsistent with the use of the primary assay in the database, 4 errors represent assay values that could not be followed up with certificates. An additional three entries were not made that affect the value used for resource estimates. There was one typographic entry error, one unknown error, and one duplicate entry, which then caused an out of sequence error. The above errors have been changed in the database.

The total number of errors found in the verification and cleaning process was 475. The verification process focused on the 370 holes used to calculate the resource estimate. A total of 61416 records (combined Collar, Survey, Geology, and Assay tables) exist in this subset of data, leading to an overall error rate of 0.7%.

14.2 Drill-hole Collar Location

Due to degradation by mining activity, reclamation, and other ground disturbance, physical location of pre-Atna drill-hole collars is impractical, if not impossible, for most ground surrounding the pit areas. However, Xxxxxx Mining Company routinely surveyed drill-hole collar locations. This was done by the mine surveyor using a total station (X. Xxxxxx, personal communication, 2004). Records for each hole were meticulously kept in a log book listing hole number, coordinate values, and elevations. This book is kept in the file system onsite. All coordinate values were obtained in local mine grid values.

The mine grid origin of 10,000N, 10,000E was established on the common corner of sections 28, 29, 32, and 33 of T48N, R24E, MDBM, now mined out. An Array of control points has been established throughout the property for survey control and has been well documented. Many of these points are still well marked and can be used for future surveys.

Homestake Mining Company drill holes were also surveyed in. These were done with GPS units and a local base station. GPS used was a Trimble LS5000 survey grade RTK system. Prior to surveying the holes a transformation from latitude-longitude values to local mine grid was established in the GPS data collector using pre-existing survey points. Mine grid coordinates were able to be read directly from the GPS unit and stored in files. (X. Xxxxxx, personal communication, 2004)

Prior to Atna’s drilling campaign, a survey was completed to see if non-survey-grade GPS would temporarily locate collar coordinates for plotting purposes. The GPS used was a Trimble AG132 with autonomous precision of 5 meters. Local grid coordinates were obtained through ArcPad by using a datum definition defined for the Xxxxxx Grid system. Settings were established in the AG 132 to read differential GPS fixes from the Coast Guard beacon system if they were available. This correction, although not nearly optimum, brought relative accuracy of the surveys to within 15 feet, acceptable for initial plotting. After system setup, the unit was taken to the field and used to resurvey control points and several (15) drill holes (water monitor xxxxx and other). Results from this survey show the GPS unit to be accurate to within 8 feet in the X direction and 6 feet in the Y direction from any survey point or known drill hole.

For underground mining, the GPS coordinate accuracy is insufficient. Due to this fact, a professional surveyor using a total station obtained Atna drill-hole collar coordinates. The survey was completed in December, near the end of drilling activity and after completion of all drill-hole pre-collars. The surveyor used an Atna geologist as xxx xxx because his knowledge of the physical drill-hole collar locations aided in the speed and accuracy of the survey. Coordinates and elevations are stored in the drill-hole collar table of the Atna database, and shown in Appendix 14-2, Drill-hole Collar Table.

15.0 Adjacent Properties

15.1 Summary

The valley east of the Xxxxxx Mountains is a discrete unnamed valley within the Humboldt River drainage basin, hosting on its west flank a series significant gold deposits. From south to north, these are:

·   Xxxxxx

·   Xxxxxx (A Zone; B Zone; C Zone; CX Zone, CX West Zone; MAG Zone; Range Front Zone)

·   Xxxxxxxx / Turquoise Ridge complex

·   Twin Creeks (an administrative consolidation of the former Chimney Creek and Rabbit Creek properties).

Deposits closest to the Xxxxxx Mountain front (Xxxxxx, Xxxxxx & Xxxxxxxx) are generally considered to be dominated by structural factors produced by high-angle faulting, while deposits somewhat distant from the mountain front (Turquoise Ridge and Twin Creeks) are credited with mineralization dominated by favorable stratigraphy-however structures play an important role in focusing these gold systems into the favorable stratigraphy.

15.2 Xxxxxx Mine

The Xxxxxx Deposit is located in the low hills east of the Xxxxxx Mountains and north of the Humboldt River, in T. 36 N., R. 41 W. (Xxxxx and Xxxxxxx, 1999). Its discovery in the 1970s was attributed to regional sampling (Xxxxxxxxxx, 1983) supported by follow-up rotary drilling. Production was deferred until 1984, when gold prices then supported mining a resource of 1.7 M tons @ 0.062 opt Au (2.1 g/t) (XxXxxxxxx et al., 2000).

The locus of this mineralization is a broad, NE-striking shear zone dipping shallowly SE, paralleling bedding. The principal host rocks are isoclinally folded carbonaceous and calcareous shales and silty limestones of the Middle Xxxxxx Formation. Silicification is the most prevalent form of alteration, and is expressed in a range of settings, from narrowly fracture-controlled, to widely dispersed in permeable horizons. Fractured settings provided higher gold grades. Multistage vein/alteration systems were recognized at Xxxxxx.

15.3. Xxxxxxxx/Turquoise Ridge Mine Complex

15.3.1 History

Historically, the focus of exploration at Xxxxxxxx was on its oxide potential, dating from the first discoveries in 1934 along the Xxxxxx Mountain front. The setting for these discoveries was fracture-controlled mineralization in the Xxxxxxxx Fault, the major N-S striking, Basin-Range fault. Deeper exploration on the fault zone resulted in high-grade discoveries.

Additional open pit resources were discovered about 3,000 feet east of the main fault zone, in an area later named Turquoise Ridge, and this area was mined out between 1993 and 1995. Some of the deeper development holes for the pit bottomed in pillow basalts believed to be unproductive. Geologists suspected that the already-mined pits might actually represent leakage above a productive system, and began a delineation program in 1994 that resulted in a commitment to shaft sinking. The shaft itself traversed a high-grade zone, and became the focus of early mine development. By the end of 1998 the production shaft was in service.

15.3.2 Geologic Setting

The Xxxxxxxx deposits are sediment hosted, micron-gold deposits whose grade is influenced by the convergence of favorable stratigraphy - often carbonaceous mudstone and silty shales, or carbonate debris flow sequences - and both high- and low-angle structures of varying orientations.

The usual formations mineralized are the Cambrian Xxxxxx, Ordovician Comus, and Ordovician Valmy. The Xxxxxx is part of the eastern platform sequence, and overlain by Cambro-Ordovician platform rocks, which include Comus carbonates. Further west, the Ordovician Comus includes both calcareous and argillaceous sediments, and interbedded mafic volcanics and tuffs. Still further west, the Ordovician Valmy(?) is represented by a more siliceous, cherty or volcanic assemblage, which is usually interpreted to be in east-directed overthrust contact with Comus beneath it. Locally, a depositional contact between Comus and Valmy is sometimes inferred. The lower Paleozoic section is usually in unconformable or high-angle fault contact with overlying Pennsylvania-Permian Xxxxxxx limestone, and the entire area was covered at one time by unconformable sheets of Miocene volcanics, now only preserved in isolated pockets.

Unraveling the true geologic relationships is significantly complicated by the overprinting effects of 4 discrete structural events:

·   The Devono-Mississippian Antler orogeny, which thrust western-facies siliceous rocks over eastern-facies carbonates;

·   The Permo-Triassic Sonoma orogeny which reactivated thrusting

·   The Jurassic-Cretaceous Xxxxxx Orogeny resulting from Pacific Plate collision North American plate;

·   The mid-Tertiary inception of regional extensional faulting.

15.3.3 Gold geology

Gold is mined underground at Placer Dome’s Xxxxxxxx Mine from two principal ore bodies; the 194 ore body at the Xxxxxxxx underground operations, and the Turquoise Ridge ore body, accessed by the Turquoise Ridge Shaft. Both ore bodies are localized at the intersections of NNS striking faults with NE striking fault zones.

Xxxxxxxx geologists have separated the local stratigraphic section into 4 units, from top to bottom (Xxxxxxx, 2004):

Unit 1: siliceous shale and greenstone;

Unit 2: a debris flow package consisting of limestone clasts in a shale and micrite matrix;

Unit 3: primarily planar-laminated calcareous siltstones and shales with limestone interbeds;

Unit 4: dominantly black mudstone and micrite.

Xxxxxxxx Underground mineralization is confined to the fault plane of the Xxxxxxxx Fault, and to calcareous/carbonaceous rocks lying within 1,000 feet of the fault. The Xxxxxxxx fault is a N-S striking 45-55 degree, east-dipping structure with several thousand feet of offset, and the 194 orebody lies in the footwall of the Xxxxxxxx Fault, hosted in units 3 and 4.

Turquoise Ridge mineralization lies several thousand feet east, in the hanging wall of the Xxxxxxxx Fault. Gold is preferentially found in horizons of carbonate debris, hosted by units 2 and 3. These units appear to be packages of debris-flow material that slumped from a carbonate platform into deeper water. The packages are volumetrically significant - the horizontal and vertical dimensions are measured in thousands of feet - and fed by a variety of faults.

Both ore bodies contain a series of dike and sill like intrusions that occur within the major fault zones. These are primarily dacite (?) in composition, but are thoroughly altered. A major difference between the 194 and Turquoise Ridge ore bodies is the presence of calc-silicate alteration at the 194 ore body; due in large part to the proximity of the Cretaceous Xxxxxx Granodiorite.

Ground conditions at the two sites differ considerably. Calc silicate alteration and hornfels, produced by the Xxxxxx stock, produced a much more competent rock at Xxxxxxxx, creating more localized fracture sets and brecciation in the host rock. This is evident in the presence of consistent bedding planes and sharp fault contacts with minimal brecciation. Rock bolts, steel sets, and steel mesh are the primary control for rock stability. Shotcrete is rarely used (<10% of drifting) in the drifts at the 194 workings.

In contrast, at Turquoise Ridge all rocks are extremely broken up and shattered, then completely argillized such that no competency exists in the host rocks at all. Underground workings are held together through bolting, mesh, and shotcrete. Xxxxxxxx geologists attribute the poor competency to a much stronger hydrothermal system at Turquoise Ridge than at Xxxxxxxx; however, the structural ground preparation and distance from the stock may be important factors as well. Turquoise Ridge is located at the intersection of NS faults parallel to the Xxxxxxxx Fault Zone and a large ENE trending fault zone extending up to Newmont Gold’s Twin Creeks Mine. These intersecting zones have produced a large area of disruption where original bedding has been destroyed by brecciation, or so badly fractured that it has no mechanical competency. The hydrothermal system then turned the rocks into clays, further destroying mechanical properties of the rock package.

15.4 Twin Creeks

15.4.1 Physiography

The Twin Creeks property lies in an alluviated valley southeast of the Dry Hills, a northeastern outlier of the Xxxxxx Range in northern Humboldt County, Nevada. The Dry Hills is an area of gently north-dipping Upper Paleozoic carbonate sediments, stratigraphically higher than rocks in the rest of the Xxxxxx Range, which consists mainly of Lower Paleozoic clastic sediments intruded by a large Cretaceous pluton, which forms the backbone of the Range.

Topography descends to the south and east from a high point in the Dry Hills (northwest of Twin Creeks), with a typical gradient of 100 feet per mile. A blanket of matrix-supported alluvium also originates to the northwest, and thickens southeastward more rapidly than the regional topographic slope, reaching thicknesses of over 700 feet three miles to the south.

15.4.2 Stratigraphy

Besides the alluvial cover, the rock sequence at Twin Creeks consists of four main elements (MacKerrow et al., 1997):

·   A group of Penn-Permian Xxxxxxx Formation mixed carbonate lithologies in depositional contact with:

·   A tectonic unit (Leviathan Thrust Sheet) of possible Devono-Mississippian age, thrust eastward over:

·   A tightly folded presumed Upper Ordovician group of predominantly basaltic xxxxx and flows (with minor interbedded shales and tuffs), and:

·   An equally tightly folded presumed Lower Ordovician group of predominantly shaley rocks (with minor mafic and basaltic igneous rocks).

The Penn-Permian Xxxxxxx Formation was divided functionally into three Members, and is the principal host in Vista Pit, with occasional extensions into northern Mega Pit. The formation is a transgressive sequence, showing a basal Lower Member of siliciclastic-dominated carbonate, a medial Middle Member dominated by massive biomicritic limestones and dolomites, and an Upper Member composed of siliciclastic sand-and-siltstones.

The Devono-Mississippian Leviathan Thrust Sheet consists of an informal Upper Member comprised mostly of exceptionally thick sequences of pillow basalts, and an informal Lower Member dominated by interbedded greenstones, cherts & tuffs. The Lower Member of the sheet crops out along the west wall of Mega Pit, where it dips about 20 degrees to the west, and sporadically appears as an east-dipping remnant along the northeast wall, suggesting that the Leviathan Thrust surface defined a north-trending antiform now cored by Mega Pit.

The Ordovician package is coherently folded into an overturned, eastward-verging Z. This package consists of three elements: a generally flat-lying, mildly arched and warped Upper Limb on the west side of Section 19 known as the Conelea Anticline, an overturned Middle Limb with no name, and a mildly arched upright Lower Limb on the east side of Section 19 known as the Tapper Anticline.

15.4.3 Structure

The axial trace of the Z-fold strikes north-northwest, and the fold plunges gently NNW at about 20 degrees. Because of the interplay of this gentle northward plunge with the regional southward slope of the bedrock surface, the practical result is that, sub-horizontal beds of the Upper Limb of the Z-fold crop out immediately beneath alluvium throughout North Mega Pit and on the western side of South Mega Pit, while the Upper Limb has been removed by erosion in the south-central part of Mega Pit, allowing the steeply dipping beds of the Middle Limb to subcrop directly beneath allluvium. Rocks of the third, Lower Limb either subcrop east of Mega Pit proper, or lie at depths of decreasing economic interest (South and North Mega Pit, respectively).

Given the size of the altered and mineralized area and the complexity of the folding, not many significant high angle faults have been recognized. The most important, from the standpoint of structural understanding, are, from oldest to youngest:

·   A district-scale Mega Pit thrust (Ear Thrust) following the Upper Axial Plane of the Z-fold in South Mega Pit, with relatively little measurable displacement.

·   A district-scale Mega Pit thrust (DZ Fault) cutting obliquely across the axial planes of the Z-fold, from the Upper Axial Plane in Mid-Mega Pit to Lower in North Mega Pit, with perhaps several thousand feet of stratigraphic throw.

·   A regional NE-trending, high angle, wide sheared zone in North Mega Pit (TC Fault) with nominal right-lateral (+/- 400') and little (50') down-to-the-north movement.

·   A family of NE-trending, high-angle faults with negligible right-lateral movement in South Mega Pit (Wry-Tail, Bill’s, and #601 Fault).

·   A long, N-S, high-angle, down-to-the-east east-dipping Mega Pit fault (CP) with differential bedrock and alluvial displacements (100 feet and 50 feet, respectively);

·   A district-scale, east-directed overthrust (Leviathan);

·   A long, N-S high-angle, down-to-the east east-dipping fault (20,000) bordering the Vista Pit area with at least 1,000 feet of displacement

As far as is known, most of the faults in Mega Pit can be shown to be pre-mineral in that they limit or compartmentalize (hydrothermal) fluid flow within the mineralized system. More often than not, only the hangingwalls of low-angle faults are mineralized, while the footwalls of high-angle faults are mineralized.

15.4.4 Erosion & Alluviation

A strongly aligned (NNW-SSE), incised erosional channel evolved along the Mega Pit corridor, on top of the originally-exposed bedrock surface, beginning somewhere in mid-Mega Pit. The erosion which centered on this drainage axis was driven in part by the extensive regional N-S alteration present in both the Upper and Lower Limb Ordovician rocks (fabrics permitting easy erosion), and the specific geometry of upturned altered Middle Limb beds, which promoted easy channel incision. This early incision eroded headward (northward) along strike, through the rest of the Middle Limb sediments, into the relatively flat beds of the Upper Limb at the northern part of Mega Pit. This channel thalweg became the depositional center of a thick column of alluvial fill, which acted as a giant reservoir of oxygenated ground water, allowing even deeper oxidation to develop. The practical result is that a large amount of Mega Pit ore was economic from the first deep excavations because the oxidation had materially lowered the required economic grade cutoffs, and allowed extraction of mineralized material that never would have been economic as sulfide ore.

15.4.5 Mineralization

Mineralization in Mega Pit occurs in several environments, including fracture fillings in what are likely feeder zones; as passive disseminations in permeable beds bounded by mafic xxxxx; and as massive replacements in structurally incompetent fold hinges. The mineral suite consists principally of auriferous pyrite, supplemented by xxxxxxxx pyrite, orpiment, stibnite, and rare realgar.

The fracture fillings require brecciation and a permeable matrix. The passive disseminations depend on prior decalcification of the sedimentary rock within the igneous 'sandwich'. The massive fold-hinge replacements depend primarily on pre-existing silicification, which can be shattered upon folding.

In spite of the variety of ore environments, the hydrothermal system is quite linear, and confined to a N-S corridor centered on the midline of Sections 6, 7, 18, 19 & 30. Along this corridor, it is clear that the mineralizing system is opportunistically exploiting all potential hosts along its longitudinal path, and. transgressing north-dipping stratigraphy from south to north, within a fairly restricted vertical interval, so that the principal hosts migrate upsection as follows:

Section 30: Middle Limb of Z-fold in Ordovician rocks

Section 19: Middle & Upper Limbs of Z-fold in Ordovician rocks

Section 18: Upper Limb of Z-fold in Ordovician rocks, & Leviathan

Section 7: Leviathan and Lower Member of the Xxxxxxx

Section 6: Lower & Middle Member of the Xxxxxxx

15.4.6 Alteration

There are seven recognizable types of alteration present at Twin Creeks. In apparent chronological order, from oldest to youngest, they are:

·   silicification I (Mega Pit)

·   propylitization (Mega Pit and Vista Pit)

·   pyritization (Mega Pit)

·   decalcification (Mega Pit and Vista Pit)

·   argillization (Mega Pit and Vista Pit)

·   silicification II (Vista Pit)

·   oxidation (Mega Pit and Vista Pit)

Silicification I appears to be early, and related almost entirely to the emplacement of the Ordovician xxxxx. This is based on the cross-sectional observation that nearly all xxxxx of any consequence are rimmed by an envelope of silicification (or extended by a layer of silicification); the envelope itself is displaced from the contact by about 3'-7'. The intervening space, between the sill and the silica envelope, is pristine because the xxxxx routinely do not hornfels their sedimentary contacts.

Leviathan Formation is universally propylitized through both pit areas. The usual characteristics are pervasive chloritic green color, calcite veining, and minor epidote.

The role and importance of pyritization at Mega Pit is indeterminate. A basic assumption is that much pyrite has been either added to, or remobilized within, the rocks. Black shales show a much higher concentration of finely disseminated pyrite than would be expected from such organically-poor rocks. Likewise, mafic phases in the basalts are often supplemented with euhedral pyrite, either as disseminations in the matrix, or as networks of hairline fractures. The role of this pyrite as a gold-enhancing agent is uncertain.

Decalcification is routinely observed in altered rock sequences whose precursors were calcareous. Effects range from slight reduction in reaction to acid to karsts and solution breccias. The decalcification has made the calcareous clastics particularly receptive to gold mineralization.

Argillization refers especially to the alteration-related development of clay fabrics, especially in the contacts of xxxxx. This appears to result from pyrite-generated sulfuric acid acting on the feldspars contained in the xxxxx. However, thinner basalts in particularly well-mineralized areas have been completely penetrated by hydrothermal fluid, allowing an extensive amount of sulfuric acid to evolve and decimate the matrix of the basalt itself.

Various styles of silicification II are recognized in the Vista Pit area. Some are associated with base-metal veins related to the Xxxxxx stock; some are strata-bound occurrences of jasperoid in the hanging-wall of the Leviathan thrust complex; and some are stockworks in greenstones.

The Leviathan thrust surface seems to control the redox boundaries throughout the mine area. In Vista Pit, the contact of Xxxxxxx Formation with Upper Leviathan defines the regional redox surface. In Mega Pit, the oxidation profile also follows the sole of the Leviathan Thrust across most of section 19, except in the vicinity of the central Middle Limb area, where the redox surface dives deeply in each of the steeply dipping shale beds between basalt xxxxx. In general, the redox surface in the horizontally-bedded areas lies at 350 feet below surface, diving to up to 1,700 feet in the overturned beds.

15.4.7 Summary

Twin Creeks is a package of intercalated impermeable basalt and permeable, mildly calcareous clastic sediments. Although the stratigraphic section dips gently north, and is folded into an overturned chevron fold, the hydrologic connectedness of the entire package allows a subhorizontal corridor of hydrothermal water access to all the permeable horizons, decalcifying the calcareous members and mineralizing them by replacement of the dissolved fabric. The corridor is surrounded to the west, east and below by a prominent front of remobilized calcite; the front above the corridor is absent by erosion. The widespread silicification is pre-mineral, and is only a contact metamorphic effect of the Ordovician intrusion of the basaltic xxxxx, and is therefore not useful for exploration.

The root zone or feeder for this system was a vertical, N-S fracture zone occurring deep in, and at the south end of, known economic mineralization, and its existence was not suspected until after years of drilling and mining. Individual high-grade accumulations occur along a rigid, N-S alignment with no visible fracture expression.

The economic value of Twin Creeks arises because the corridor of altered rock lies in a downfaulted block in an extension terrain. The vertical extent of alteration allowed deep erosion, and deposition of a thick cap of alluvium over the buried deposit. Highly oxygenated water in the alluvial cap promoted deep oxidation of the bedrock, converting marginal gold sulfide grades to economic oxide grades.

16.0 Mineral Processing and Metallurgical Testing

This technical report is in support of a resource calculation only. While it is anticipated that the project’s mineralized material will have metallurgical and mineral processing characteristics similar to those at the adjacent, producing mine properties of Xxxxxxxx and Twin Creeks, no detailed metallurgical studies have been completed by Atna Resources on the resource zones summarized in this technical report.

17.0 Mineral Resources Estimates

17.1 Introduction

The mineral resource estimates for the CX and Range Front zones were made from 3-dimensional block models generated using commercial mine planning software (MineSight).

The project limits are based on a local Xxxxxx mine grid system initially developed during mine operations some years back The mine grid is not rotated and the origin (point 10000,10000) is located at 478294.7E, 4553517.9N in the UTM Zone 11, North American Datum of 1927 coordinate system. The mineralized zones occur roughly between 8500E to 11000E, 9500N to 13500N and 2500 ft to 5500 ft elevation. The project limits are described in more detail in section 17.9.

A series of vertical cross sections, oriented approximately perpendicular to the general trend of the mineralized zones, are utilized for viewing the deposits in cross section. This series of 40 sections, numbered 5300NE through 9300NE, are oriented at an azimuth of 120 degrees and are spaced at 100 ft intervals as shown in Figure 17-1.

Figure 17-1: Rotated Vertical NW-SE Cross-Sections

The mineral resource estimate was generated from drill-hole sample assay results and a geologic model which relates to the spatial distribution of gold. Individual domains, reflecting distinct zones or types of mineralization, have been defined and interpolation characteristics have been defined for each domain based on the geology, drill-hole spacing and geostatistical analysis of the data. The degree of confidence in the resources have been classified based on the proximity to sample locations and/or surface exposures and are reported, as required by NI 43-101, according to the CIM standards on Mineral Resources and Reserves.

This report includes estimates for mineral resources. There are no mineral reserves prepared or reported in this Technical Report.

17.2 Geologic Model, Domains and Coding

17.2.1 Geologic Model

The CX and Range Front are two sub-parallel fault zones which are separated by approximately 600 feet and dip to the southeast at about 55 degrees. These fault zones range in thickness from 30 to over 300 feet and average approximately 150 feet. The majority of the fault is relatively barren of mineralization; however, it does host a distinct, somewhat dendritic web or zone that is highly enriched in gold. The nature and distribution of the mineralization is thought to result from solutions which passed through the “path(s) of least resistance” depositing microscopic gold in the receptive host.

Grade estimations conducted without a distinct internal HG domain boundary would result in the averaging of high and low grade samples throughout the fault zone resulting in a large but low-grade model. This contradicts the nature and distribution of the gold seen in both open pit exposures and in the drill holes.

The enriched, or high-grade (HG), portion of the fault zone is often defined by the presence of silica (veins), pyrite, or the visual identification of traces of orpiment, realgar, stibnite and/or cinnabar. Unfortunately, the various vintages of drilling on the property has resulted in a database which, in areas, lacks the details such as the identification of these trace minerals. As a result, it is difficult to produce a geologic model of the HG portion of the fault zone based on the available geological information.

Therefore, in order to reproduce the natural distribution of the mineralization within the fault zones, a separate HG zone domain has been developed through a combination of geostatistical applications and geologic interpretation of the sample data. This approach has resulted in a grade model which closely reflects the trends observed in the surface exposures and the drilling results.

17.2.2 Domains and Coding

The geologic model for the Xxxxxx gold-bearing zones has been developed in two stages. Initially, the CX and Range Front fault zones have been developed based on the geologic logging information. Second, the high-grade portions within the fault zones are defined through a combination of geostatistical and manual methods.

The CX and Range Front fault zones have been identified in drill holes based on information defined in the drill logs. This process included a review of core photos (where available) and re-logging of stored core/cuttings from historic drill holes where required.

The fault zone intervals were displayed along the drill-hole traces and the subsequent geologic interpretation based on this data was conducted on cross sections. The interpreted sectional polylines were then linked to produce 3-dimensional wireframe solids. The CX and Range Front fault zones are shown in cross section in Figure 17-2 and in an isometric view in Figure 17-3.

It should be noted that geologic model has been extended above the current surface topography in the area of the CX pit. The information gained from including the mined-out portion of this zone enhances the estimation of the remaining CX resource. The final contained resource in this zone is limited to the current topographic surface.

Figure 17-2: Cross Sectional View of CX and Range Front Fault Zones

Figure 17-3: Isometric View of CX and Range Front Fault Zones

Following the generation of the fault zones, the high-grade gold-bearing domains (HG zone) were defined as follows. The drill-hole samples were composited to standard 5-foot sample lengths (this step is described in more detail in section 17.4). Samples contained within the fault zones were evaluated both visually and statistically (described in detail in section 17.5) and an indicator threshold of 0.1 opt Au was selected to define the HG portion of the fault zone. Samples within each of the faults were assigned a value of zero (0) if the grades were below 0.1 opt Au and a value of one (1) if the grade was above 0.1 opt Au. Probability estimations were then conducted defining the likelihood of the presence of the HG zone located within the fault zone boundaries. The indicator probabilities within blocks are estimated using an inverse distance weighting (IDW) technique with search directions controlled by “relative elevations” as described below.

Although the CX and Range Front fault zones occur as relatively planar structures, local undulations exist as shown in figures 17-2 and 17-3. Based on drilling data and surface exposures, the HG zones located within the host fault zones show undulations and local branching, or bifurcation, into multiple zones. The indicator probability estimates using regular ellipse-oriented IDW or ordinary kriging (OK) methods resulted in very planar distributions which are not representative of the true nature of the HG zones. The “relative elevation” (RE) approach assigns temporary elevation values to both drill-hole composites and the model blocks, relative to the location with respect to a 3-D surface. Then, during the resulting estimation, the search orientation is controlled by the shape of the designated surface. Note that similar results could be achieved by splitting the deposit into a number of individual domains, each with a unique search ellipse or set of variogram parameters. However, this would be a difficult and potentially inaccurate approach due to the relative lack of sample data over some portions of the deposit.

The controlling surfaces for the RE estimation were initially interpreted on cross sections defining the general trend of the HG sample intervals with respect to the overall shape of the host fault zones (Figure 17-4). The interpreted lines were joined to form 3D surfaces, one for the CX zone and one for the Range Front zone. The relative elevations with respect to these surfaces were then assigned to the composite samples and model blocks and the indicator probabilities were estimated for each of the two zones. Figure 17-5 shows a comparison of the indicator probability distributions resulting from the relative elevation method versus a regular ellipse oriented IDW estimate (note that the probability distribution resulting from an OK application produces very similar results to the regular ellipse oriented IDW estimate). Figure 17-6 shows a similar comparison, in cross section, from the Range Front zone.

Figure 17-4: Sectional View of HG Zone Trends Within Fault Zones

Figure 17-5: Comparison of Indicator Probability Estimation Techniques - Plan

Figure 17-6: Comparison of Indicator Probability Estimation Techniques - Section

A series of probability grade shells were produced from the block indicator probability estimates. A “grade shell” is essentially a 3-dimensional contoured surface which envelops volumes at defined threshold limits. Probability grade shells were produced from 25% to 50% probability limits at 5% increments. The probability shells were then compared to the original gold assay data on section and plan. In general, the probability shells performed well in areas with relatively high sample density. However, in areas where the drill-hole spacing increases (typically at depth) or where the HG zone became quite narrow, the probability shells did not consistently represent the continuity of the zone. Ultimately, the 30% probability shell for the CX zone and the 35% shell for the Range Front zone were selected as the best fit with the sample data.

Some minor modifications were made to the probability shells as follows. The lower portion of the CX zone shell, which is not bounded by additional drilling, was limited to a distance of 170 feet from a drill hole (Figure 17-7). Also, a thin zone on the north part of the shell was manually added to more accurately represent the continuity encountered in the drilling in this area.

Figure 17-7: CX High-Grade Zone Modifications

Several additions were made to the Range Front probability grade shell in areas where the indicator approach was not able to reproduce the continuity of the narrow portions of the deposit. These are shown in Figure 17-7. As in the CX zone, the lateral extent of the grade shell in mineralized areas not bounded by additional drilling, was limited to a distance of 150 feet from a drill hole.

Figure 17-8: Range Front HG Zone Modifications

The geologic domains are summarized in Table 17-1.

Table 17-1: Summary of Geologic Domains

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

17.3 Available Data

The exploration and production history on the Xxxxxx property dates back over more than 50 years and as a result, there is a rather extensive drill-hole database consisting of almost 3000 holes. The resource estimate described in this report is based on the information provided from a total of 401 drill holes which are in close enough proximity to the CX and Range Front zones to have an influence on the resource modeling.

Atna Resources drilled a total of 31 holes during its 2004/05 program. As of the data freeze date for this resource estimate (February 15, 2005), the assay results from holes APRF-229A and APRF-230 have not been received, however, the geology logging information was utilized in the development of the Range Front fault zone domain.

The total length of drilling in the 401 holes is 236,219 feet, of which 52 holes are diamond drill (DD) cored holes and the remaining 349 are reverse circulation (RC) holes. Note that in an effort to reduce costs, many of the recent Atna holes were pre-collared and cased using a reverse circulation drill rig with a conversion to DD prior to intersecting the target fault zone. Although RC drilling comprises approximately 85% of the holes in the database, the majority of this type of drilling is limited to the shallow upper-portion of the deposits as shown in Figures 17-9 and 17-10. As a result, the diamond drilling results contribute more in relation to the overall resource estimation. It is estimated that DD holes cover approximately 65% of the lateral extent of the CX zone and approximately 85% of the area of the Range Front zone.

Figure 17-9: Distribution of Drill Holes by Type - CX Zone

Figure 17-10: Distribution of Drill Holes by Type - Range Front Zone

Samples in the database which were analyzed but returned values below the detection limit were assigned values equal to one half the detection limit.

There are a total of 45,121 samples analyzed for gold in the database. The majority (89%) of the sample intervals are a standard length of 5 feet; however they range from a minimum of 0.2 feet to a maximum of 82 feet. These large intervals are composite samples collected from primarily rotary holes in areas where gold mineralization is not anticipated.

The presence and location of the CX and Range Front fault zones has been identified from drill log data, core photos or re-logging of stored core, in a total of 209 holes in the database. The fault domains have been developed from the information provided by these drill holes. The remaining 192 holes in the database intersect one of the fault zones but the actual location of the fault in the hole cannot be confirmed (i.e. It is not clearly identified in the drill logs, there are no photos of the core and there is no material stored from these holes).

17.4 Compositing

Compositing of drill-hole samples is performed in order to standardize the database for further statistical evaluation. This step eliminates any bias related to the sample length that may exist in the data.

As stated previously, the vast majority (89%) of the original sample intervals are 5 feet in length. As a result of this fact, the composite length was also selected at 5 feet. The reason for this is that the underlying nature of the data is retained in the composited data, where the selection of a longer composite length introduces some smoothing, which may mask some of the natural features of the data.

Drill-hole composites are length-weighted and have been generated “down-the-hole” meaning that composites begin at the top of each hole and are generated at 5-foot intervals down the length of the hole.

Prior to compositing, the original drill-hole samples were “speared” with the fault ZONE domain solids. This step tags each sample interval with the appropriate ZONE-code designation (ZONE=1 for the CX fault and ZONE=2 for the Range Front fault zone). These contacts are then honored during compositing (i.e. composites begin and end at the ZONE domain boundaries).

Several holes were randomly selected and the composited values were checked for accuracy. No errors were found.

17.5 Exploratory Data Analysis

Exploratory data analysis (EDA) involves the statistical evaluation of the database in order to quantify the characteristics of the data. One of the main purposes of this exercise is to determine if there is evidence of spatial bias which may require the separation and isolation of domains during interpolation. The application of separate domains prevents unwanted mixing of data during interpolation and the resulting grade model will better reflect the unique properties of the deposit. However, use of domain boundaries in areas where the data is not statistically unique may impose a bias in the distribution of grades in the model.

A domain boundary, which segregates the data during interpolation, is typically applied if the average grade in one domain is significantly different from that of another domain. A boundary may also be applied where there is evidence that there is a significant change in the grade distribution across the contact.

The various domains that reflect differing styles and distribution of gold mineralization in the deposit are listed in Table 17-1. The drill-hole composites have been speared with these domains prior to statistical analysis. EDA evaluation of these domains includes basic statistics, contact profiles, histograms and log-probability plots.

17.5.1 Basic Statistics by Domain

The basis statistics of samples contained within the CX and Range Front fault zones are summarized in Table 17-2.

Table 17-2: Drill-hole Sample Statistics by Fault Zone

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

(Statistics based on 5-foot drill-hole composite samples, weighted by sample length)

Table 17-3: Drill-hole Sample Statistics by HG/LG Portion of Fault Zone

Gold (opt)	CX HG Zone	CX LG Zone	RF HG Zone	RF LG Zone
Total #/ft samples	564/2,819	6,186/30,842	248/1,228	2,415/12,041
Min	0.0005	0.0001	0.0009	0.0001
Max	1.840	1.305	2.337	1.054
Mean	0.241	0.011	0.295	0.011
Std Dev	0.220	0.031	0.277	0.035

(Statistics based on 5-foot drill-hole composite samples, weighted by sample length)

The results from the tables above indicate that the Range Front zone is slightly higher grade than the CX zone. Although the physical extent of the Range Front zone is larger, the CX zone has more contained samples due its higher density of drill holes.

Note that the low-grade (LG) portion of the fault zones contains several relatively high-grade samples (they are described in more detail in section 17.7 of this report). These were considered somewhat anomalous during the development of the HG zone domains because they could not be included in the interpretation with any degree of confidence. Therefore, at present, they remain outside of the HG zone domain. The understanding of these intervals is expected to improve with additional drilling and it is anticipated that they may add to overall resource in the future.

17.5.2 Contact Profiles

The nature of grade trends between two adjacent domains is evaluated using the contact profile which graphically displays the average copper grades at increasing distances from the contact boundary. Contact profiles that show a marked difference in grade across a domain boundary, are an indication that the two data sets should be isolated during interpolation. Conversely, if there is a more gradual change in grade across a contact, the introduction of “hard” limitations to the data (i.e. segregation during interpolation) may result in very different trends in the grade model - in this case the change in grade between domains in the model is often more abrupt than the trends seen in the raw data. Finally, a flat contact profile indicates no grade changes across the boundary. Here, the segregation of the data, or the lack of separation (i.e. “hard” vs. “soft” domain boundaries) will produce similar results in the model.

The very nature of the construction of the CX and Range Front HG zones produces domains which differ in grade from the surrounding host. The contact profile shown in Figures 17-11 and 17-12 clearly show the change in gold grade across the HG zone boundaries.

Figure 17-11: Contact Profile of CX HG vs. LG Domains

Figure 17-12: Contact Profile of Range Front HG vs. LG Domains

17.5.3 Histograms and Log-Probability Plots

The sample data contained in the various domains is displayed in a series of histograms and probability plots (Figures 17-13 through 17-24).

The histograms of the fault zones as a whole show very skewed distributions of data. However, once separated, the HG vs. LG histograms are, as expected, significantly different. The log-probability plots for the fault zones (Figure 17-19 and Figure 17-22) exhibit subtle breaks in the slope of the distribution at a grade of 0.1 opt Au. This feature is more evident in the Range Front fault.

This break is accentuated when the HG and LG data is viewed separately. For example, Figure 17-20 shows the distribution of samples inside the CX HG zone. There are a series of samples below the 0.1 opt Au threshold for the HG zone that show a very different distribution trend when compared to the higher-grade samples. These low-grade samples represent narrow intervals which are bracketed by higher-grade material and are the result of generalizations made during the interpretation of the HG domain boundaries. It is expected this mixing of domains will be reduced with additional detailed drilling information.

Figure 17-13: Histogram of Gold in CX Fault Zone

Figure 17-14: Histogram of Gold in CX HG Zone

Figure 17-15: Histogram of Gold in CX LG Portion of Fault Zone

Figure 17-16: Histogram of Gold in Range Front Fault Zone

Figure 17-17: Histogram of Gold in Range Front HG Zone

Figure 17-18: Histogram of Gold in Range Front LG Portion of Fault Zone

Figure 17-19: Log-Probability Plot of Gold in CX Fault Zone

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

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

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

17.5.4 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.

17.6 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 Xxxxxx 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.

17.7 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 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.

17.8 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 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 (Xxxxxx 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 included as Appendix 17-1 (CX Correlogram) and Appendix 17-2 (Range Front Correlogram), and are summarized in Table 17-5.

17.9 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 17-23.

Table 17-5: Variogram Parameters

Domain	Nugget	S1	S2	1st Structure			2nd Structure
				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

Figure 17-25: CX Zone Block Model Limits

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

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 17-24 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)

17.10 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.

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

17.11 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.

17.11.1 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.

17.11.2 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 Xxxxxxx 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 foe the CX and Range Front HG zones in Figures 17-27 and 17-28.

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

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

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.

17.11.3 Comparison of Interpolation Methods

The OK, IDW and NN models are tabulated for comparison purposes in Table 17-10.

Table 17-10: Comparison of Interpolation Methods

Cut-off	OK Model		IDW Model		NN Model
(Auopt)	kTons	Auopt	kTons	Auopt	kTons	Auopt
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.

17.11.4 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 17-29, 17-30 and 17-31 and the Range Front HG zone shown in Figures 17-32, 17-33 and 17-35. 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.

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

 

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

 

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

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

 

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

 

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.

17.12 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 X. Xxxxx, 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).

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 17-5). 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 17-11 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%

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 17-35). 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 Xxxxxx deposits are geologically similar to others in the area and the mode of formation is well understood. The nearby Xxxxxxxx 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 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.

Figure 17-38: CX Zone Resource Classification

 

Figure 17-39: Range Front Zone Resource Classification

17.13 Key Assumptions of Economic and Technical Parameters

Generalized economic and technical parameters have been estimated for the Xxxxxx project in order to project the relative scale of the potential economic viability of the 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 Xxxxxx deposits.

The key economic and technical assumptions, parameters and methods used in this report are primarily derived from proximal existing operations in the Xxxxxx area. Many of the mines in Northern Nevada exhibit similar ore 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 be 700 tons per day.

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

5) Processed by toll milling at third-party mill (Newmont’s Twin Creeks, Gold Quarry, or Lone Tree xxxxx; Xxxxxxx’x Goldstrike complex; and/or Queenstake’s Jerritt Canyon mill). Toll milling costs are estimated to be $25/ton milled, including transportation. This is based on initial discussions between Atna and the third-party operators and assumes that 100% of Xxxxxx 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 0.20optAu in tables)

17.14 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: Mineral Resource, CX Zone

Category	Cut-off (Au opt)	Tons (000)	Grade (Auopt)	Contain 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 +	0.15	947	0.28	262
Indicated	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 = 0.20optAu)

Table 17-13: Mineral Resource, Range Front Zone

Category	Cut-off (Au opt)	Tons (000)	Grade (Auopt)	Contain 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 +	0.15	811	0.32	257
Indicated	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 = 0.20optAu)

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

Category	Cut-off (Au opt)	Tons (000)	Grade (Auopt)	Contain 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 +	0.15	1,758	0.30	519
Indicated	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 = 0.20optAu)

18.0 Other Relevant Data and Information

Two drill holes were completed after the cut-off date for data included within this Technical report. APRF-229A and APRF-230, illustrated in Figure 18-1, were both finalized after February 15^th, 2005, the effective date of this report.

Figure 18-1. Location of Final Phase I Range Front Holes (APRF-229A and APRF-230)

The assay results from these two holes, shown below in Table 18-1, are as follows:

Table 18-1: Significant Drill Results From Holes APRF-229A and APRF-230

Drill-hole ID	Total Depth	From (ft.)	To (ft.)	Intercept Length (ft.)	Intercept Length (meters)	Grade (ounce/ton) Gold	Grade (gram/tonne) Gold
APRF-229A	1,515.0	1,466.0	1,474.5	8.5	2.59	0.458	15.70
APRF-230	924.5	730.0	736.0	6.0	1.83	0.635	21.78
	including	734.0	736.0	2.0	0.61	1.108	37.97

Drill hole APRF-229A cut mineralization consistent with mineralization extrapolated from drill holes completed prior to the data cut-off date for this report and these results will not materially effect the results reported within this Technical Report. Drill hole APRF-230 was lost due to poor drilling conditions prior to fully penetrating the targeted Range Front zone. APRF-230 did cut significant mineralization well above the zone, as displayed above in Table 18-1. However, because this mineralized intercept is outside the resource evaluation envelopes, this data will have no effect on the resource calculation completed within this Technical Report.

19.0 Interpretation and Conclusions

Current data indicates the Xxxxxx property contains two major zones of gold mineralization typical of other gold deposits in the district and region. The CX and Range Front zones are major brittle structural features that focused hydrothermal fluid flow and gold mineralization. Analytical results from drilling completed through February 15, 2005 by Atna and previous operators has outlined two gold-bearing mineral zones which remain open in several areas along strike and dip. These two zones display numerous physical and geochemical similarities to the neighboring Xxxxxxxx, Turquoise Ridge, and Twin Creeks Mines, all of which are considered Xxxxxx-type, sediment-hosted gold deposits and are currently being commercially exploited. Drill-hole intercept data is sufficiently dense within the CX and Range Front zones to provide an adequate level of confidence for the resource estimates established in Section 17 of this report.

The resource modeling techniques utilized in this Technical Report are considered consistent with practices currently in use in the mining industry. The density of the current drill coverage, primarily at depth, is relatively wide-spaced and, as a result, a high proportion of the resource remains in the Inferred category. Additional drilling is required to raise confidence levels, particularly in the deeper portions of both the CX and Range Front, where drill-hole xxxxxx-points are generally on greater than 200-foot centers. Future drilling campaigns are planned with this objective in view, and should be part of an overall program to complete a pre-feasibility study, mine plan and establish mineral reserves within the CX and Range Front zones.

20.0 Recommendations

The next exploration phase at the project should focus on more fully defining the nature and distribution of gold within the Range Front and CX resource blocks. This phase should focus on the continuation of shallow Range Front surface drilling, and on development of underground access to both the Range Front and CX mineral zones.

Drilling in the shallow Range Front environment will enhance definition of newly discovered gold resources lying above 4500' elevation (see Figure 9-2). Development of underground access will provide platforms for expanded drilling of the existing resource blocks in both fault zones (see Figure 17-2) and raise the confidence level attributable to these blocks. An initial drill spacing of 100-feet by 100-feet (both surface and underground) is recommended for the shallow portion of the Range Front zone to refine the understanding of the geology, structural controls and detailed distribution of gold grades.

Detailed drilling will also refine the variography for future resource modeling and define the appropriate drill spacing for reserve development. It will also generate sufficient mineralized material for preliminary metallurgical testing and process metallurgical planning.

Once detailed drilling is completed (or substantially completed) on the upper Range Front zone, the underground workings should be extended into the mineralized zone and a bulk sample collected for detailed metallurgical testing.

Development of underground access and bulk sampling of the mineralized zone(s) will provide necessary cost and planning data for development of an economically sound mine plan allowing profitable exploitation of the gold resources outlined and discussed in this Technical Report.

After establishing an initial exploration level for resource definition drilling, the underground workings should be extended as a decline to provide access to deeper levels of the mineral zones for continued up-grading of the resource confidence levels and conversion of the “Inferred Resource” blocks into Indicated and/or Measured resource categories. A minimum drill-hole spacing of 200-foot centers is necessary to evaluate the deeper resource blocks, based upon the current data. The drilling recommended in the shallow portions of the Range Front zone will directly influence the appropriate drill-hole spacing throughout the deposit(s), and it is certain that additional work will be necessary as deeper underground access is achieved and reserve development drilling is being conducted.

All underground workings should be structurally and geologically mapped in detail - a scale of 1:240 (1 inch = 20 feet) is recommended. A face-and-rib sampling program should be maintained throughout the underground drifting program, and the collected data added to the resource modeling database.

Concurrent with the development of the underground access, several reverse circulation rotary drill holes should be “twinned” with core holes. It is recommended that at least six (6) earlier reverse circulation rotary drill holes be twinned with core holes designed to cut the mineralized zone within ten (10) feet of the existing drill-hole intercepts. Several of these holes may be completed as underground drill holes as part of the definition drill program outlined above. If the core results are comparable to the original RC sample results, this exercise will validate RC sample integrity and bolster the confidence level of these intercept values within the database.

Detailed surface mapping and sampling of the CX mineral zone is recommended within the CX pit (where pit wall and bench conditions are sufficiently safe to allow such work). Mapping at a scale of 1:240 and detailed continuous channel sampling of the bench walls perpendicular to the strike of the mineralized zone are recommended. Results from this work will then be added to the resource database and models.

Exploratory drilling along strike and down-dip of the zones is justified, as both the CX and Range Front zones remain open to expansion. While not the subject of this Technical Report, which focuses on the CX and Range Front mineralized zones, additional zones of exploitable mineralization may also be discovered in association with the numerous drill-hole gold intercepts which occur well outside the mineralized envelopes used in this study. Additional exploration drilling is warranted and recommended as work progresses on the project.

As an adjunct of the additional drilling and underground work discussed above, work should continue on elimination of database errors both within the existing data set and on all new data added to the database. Atna’s current procedures are ensuring data reliability and integrity, but continued vigilance is an integral part of making successful economic decisions as the project progresses towards exploitation.

20.1 Recommended Program and Budget

Based upon the recommendations described above, a program has been designed to focus on the Range Front zone’s mineralization between the surface and the 4,400-foot level where both thicknesses and grades best suited for the project’s initial development (if reserves are established). The following program and budget has been designed with the initial focus on the Range Front zone mineralization above the 4,400-foot level:

Surface Drilling-Range Front Zone	US$
Rotary pre-collar drilling
20 holes, 7,500 feet @ US$25/foot	$187,500
Diamond core drilling (core-tails)
20 holes, 8,500 feet @ US$50/foot	$425,000
Surface drill sample prep., assays & standards
5,000 samples @ $20/sample	$100,000
Surface roads and drill pads (dozer work)	$10,000
Underground Drilling-Range Front Zone
32 holes, 16,000 feet @ US$50/foot	$800,000
Underground drill sample prep., assays & standards
5,000 samples @ $20/sample	$100,000
Drilling subtotal	$1,622,500
Metallurgical Testing
Sample compositing, direct cyanidization, roasting, and autoclave testing	$100,000
Metallurgical subtotal	$100,000
Underground adit/drill crosscut development
Contractor mobilization	$100,000
Portal Development	$75,000
Ventilation - gen sets	$125,000
3,500 feet of drifting @ US$1,200/foot	$4,200,000
Mining subtotal	$4,500,000
Dewatering
Dewatering well development
2 - 28 inch xxxxx @ 600 feet	$1,500,000
Pumps, piping, & discharge area development	$1,000,000
Dewatering subtotal	$2,500,000
Permitting
Consulting fees, permitting, groundwater, etc	$150,000
Reclamation bonding	$500,000
Permitting subtotal	$650,000
Project Team - personnel
Project manager (engineering)	$100,000
Project geologist	$100,000
3-Support geologists	$240,000
3-geotech	$150,000
1-administrative/accounting	$60,000
Personnel subtotal	$650,000
Land costs
Land payments & claim fees and expenses	$250,000
Land costs subtotal	$250,000
Miscellaneous costs
Rentals (field vehicles)	$50,000
Field materials and supplies	$100,000
Office materials, office utilities	$50,000
Project insurance	$100,000
Computers, software, & all other	$100,000
Miscellaneous costs subtotal	$400,000
Project total	$10,672,500
Contingency @ ~12%	$1,324,500
Proposed Project Budget	$12,000,000

References

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

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Xxxxxxxxx, X., Xxxxxxxxx, X., Xxxxxxxx, M., Xxxxxx, B. and Xxxxxxx, E., 2000, Geologic overview of the Xxxxxxxx Gold Mine geology, exploration and ore deposits, Humboldt County, Nevada, in Xxxxxxxx, 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 Xxxxxxxx region, Humboldt County, Nevada, p. 113-121.

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

Xxxxx, D.A. and Xxxxxxx, X.X., 1999, Gold and silver resources in Nevada: Nev. Bureau of Mines & Geology Map 120.

Xxxxxxx, X., 2004, Internal company memoranda.

Xxxxxxxx, X.X. and Xxxxx, X.X., 1974, Geologic Map of the Iron Point Quadrangle, Humboldt County, NV: U.S. Geological Survey Geologic Quadrangle Map GQ-1175, scale 1:24,000.

Xxxxxx, X.X. and Xxxxxxxxxx, E.L., 1991, Geology of the MAG deposit, Xxxxxx Mine, Humboldt County, Nevada, in Xxxxxx, X.X., Xxxxx, R.E., Xxxxxxx, X.X., and Xxxxxxxxx, X.X. (eds.), Geology and ore deposits of the Great Basin: Geological Society of Nevada, 1990 Symposium Proceedings, p. 845-856.

Xxxxxx, X.X., 1994, Gold in xxxxxxxx framboidal pyrite in deep CX core hole DDH-1541: Unpublished Xxxxxx Mining Company Report, June 21, 1994.

Xxxx, B., Xxxxxxx, S., Xxxx, K. and Xxxxxxx, T., 2004, Xxxxxx Joint Venture Field Trip, in Xxxxxxxx, R., Xxxxxxx, E. and X'Xxxxxx, 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 Xxxxx, X.X., 2000, Characteristics and models for Xxxxxx-type gold deposits: Reviews in Economic Geology, v. 13, p. 163-220.

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

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

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XxXxxxxxx, X.X., Xxxxxxxxxxx, X.X. and X.X.Xxxxxxxx, 2000, The gold deposits of Xxxxxx Mining Company: a review of the geology and mining history through 1999, Humboldt County, Nevada, in Xxxxxxxx, 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 Xxxxxxxx region, Humboldt County, Nevada, p. 123-153.

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Xxxxxxx, X.X., 1980, Geology of Nevada, Nevada Bureau of Mines and Geology Special Publication 4, 136 x.

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Winnemucca, NV web site (xxxx://xxx.xxxxxxxxxx.xx.xx/)