CHIEF CONSOLIDATED

MINING COMPANY

THE TINTIC MINING DISTRICT

(UTAH)

PROSPECT AND RETROSPECT



ByBen Ainsworth, P. Eng.Derek Barratt, P. Eng.David Jenkins, P. Geo.Rod McElroy, Ph.D.Brian Mountford, P.
				  Eng.


NOVEMBER 2005





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TABLE OF CONTENTS



Page No.SUMMARY
			 
			 1TERMS OF
				  REFERENCE
			 
			 4PROPERTY
				  DESCRIPTION AND LOCATION
			 
			 5ACCESSIBILITY,
				  CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY
			 
			 7HISTORY
			 
			 8GENERAL
			 
			 8THE TRIXIE
				  MINE
			 
			 12BURGIN
			 
			 14GEOGRAPHICAL
				  SETTING
			 
			 20DEPOSIT
				  TYPES
			 
			 22THE NEW BURGIN
				  MINE
			 
			 24BACKGROUND
			 
			 24FUTURE
				  PROGRAM
			 
			 27MINERAL
				  PROCESSING
			 
			 29METALLURGICAL
				  TESTWORK
			 
			 33BURGIN
				  WATER
			 
			 37HYDROLOGY
			 
			 37MINE
				  DEWATERING
			 
			 39Comments
			 
			 41WATER
				  APPROPRIATION
			 
			 41DESALINATION
			 
			 43Summary
			 
			 43Background
			 
			 44Scope and Battery
				  Limits
			 
			 45Exclusions
			 
			 45Overview and
				  Technology Selection
			 
			 45Overall Treatment
				  Concept
			 
			 46Reverse Osmosis
				   Process Considerations
			 
			 46Treatment
				  Costs
			 
			 47Cost
				  Factors
			 
			 47Direct Capital
				  Costs
			 
			 48Indirect Capital
				  Costs
			 
			 48Contingency
			 
			 48Operating
				  Costs
			 
			 48Project
				  Rationale
			 
			 49All-in Cost
				  Comparators
			 
			 49Capital Cost
			 
			 50Operating
				  Cost
			 
			 50Conclusions and
				  Recommendations
			 
			 51






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TABLE OF CONTENTS(Contd)



Page No.POTENTIAL WATER
				  SALES
			 
			 52ZUMA CLAY
				  PROJECT
			 
			 54GENERAL
			 
			 54RESOURCES
			 
			 54HALLOYSITE
			 
			 55USES AND MARKET
				  POTENTIAL
			 
			 56OTHER SOURCES OF
				  HALLYOSITE
			 
			 57ZUMA SHAFT AND
				  LARSEN CLAY PITS
			 
			 58RECOMMENDED WORK
				  PROGRAM
			 
			 59EXPLORATION
				  TARGETS (OTHER THAN THE BURGIN)
			 
			 61ZONE
				  A
			 
			 61Recommendation
			 
			 65BALL PARK
				  ZONE
			 
			 67Recommendation
			 
			 72OTHER EXPLORATION
				  TARGETS
			 
			 73NEAR SURFACE
				  EXPLORATION POTENTIAL
			 
			 76District in
				  General
			 
			 76Exploration and
				  Prospectivity
			 
			 77Exploration
				  Targets
			 
			 77Trixie
				  Ore
			 
			 82Burgin
				  Tailing
			 
			 82ENVIRONMENT/PERMITTING
			 
			 83REFERENCES
			 
			 84






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TABLE OF CONTENTS(Contd)







|FIGURES AND
				TABLES |  | Page No.                                            
|---------------------- |  | ----------------------------------------------------
|                       |  |                                                     
|Figure 1               |  | Development Program and
				Budget                  
|Figure 2               |  | Property Location Map                               
|Table No.1             |  | Metal Production from the largest
				Mines         
|Figure 3               |  | Property Outline and Defined
				Targets            
|Figure 4               |  | Burgin Surface                                      
|Figure 5               |  | Burgin Extension Deposit &
				Drill Holes          
|Table No.2             |  | Historical Resource/Reserve
				Estimates           
|Table No.3             |  | Deeper Drill Hole
				Intersections                 
|Table No.4             |  | Recommended Drill Holes                             
|Figure 6               |  | Burgin Mine Metallurgical Flow
				Sheet            
|Table No.5             |  | Flotation Testwork                                  
|Table No.6             |  | Separate Pb/Zn
				Concentrates                     
|Table No.7             |  | Recommended Drill
				Program                       
|Table No.8             |  | Possible Water Solutions                            
|Table No.9             |  | Capital & Operating
				Costs                       
|Table No.10            |  | Desalination Operating
				Costs                    
|Table No.11            |  | Current Water Costs                                 
|Figure 7               |  | Exploration Targets Adjacent to the
				Burgin Mine 
|Table No.12            |  | Drill Hole Intersections from Zone
				A          
|Figure 8               |  | Exploration Targets Ball Park
				Section A  A ́    
|Figure 9               |  | Exploration Targets Ball Park
				Section B  B ́    
|Table No.13            |  | Drill Hole Intercepts  Ball
				Park Mineralization
|Table No.14            |  | Ball Park Estimates                                 
|Table No.15            |  | District Targets and
				Potential                  



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SUMMARY

Chief Consolidated Mining Co. (Chief) owns
		considerable mining claims comprising some 19,500 acres that control the Tintic
		and East Tintic Mining Districts in Utah. These claims have been subjected to
		intermittent underground mining operations from several operating mines for
		over 140 years.

The principal authors of this report, after
		spending time on the property and an extensive evaluation of the volumes of
		historic data, have concluded that significant opportunities exist for further
		development and exploitation of the claims.

A program of development is recommended, at
		a cost of $1,170,000 to complete a first phase of a plan to increase the value
		and enhance the prospects of viably operating the following four projects.(See Figure No.1)


			 Advancing the Burgin extension
				  deposit through to a bankable Feasibility Study.



			 Further development of the concept
				  to sell potable water from a desalination plant fed by pumped water from the
				  Burgin Mine.



			 Investigation of the economic
				  possibilities of producing and selling halloysite (and other) clays from the
				  Zuma and other areas.



			 Continued investigation and
				  development of other exploration targets in the claim block and adjacent
				  areas.


A success contingent, phased approach is
		recommended since, other than the Zuma clay potential, the areas of interest
		are related.

The Burgin Mine project should be moved
		forward by creating access to the 1050 level via the Apex No.2 shaft and
		drilling a series of holes into the down-dip extension of the previously mined
		ore-body, as well as satellite bodies and targets. The objectives of this work,
		at a cost of $860,000, will be to provide a N1 43-101 compliant resource and
		subsequent reserve; to obtain samples for metallurgical testing and to
		ascertain initial hydrological characteristics relative to ore deposit
		dewatering.

The Burgin Water project should be developed
		along three parallels.


			 Effective dewatering of the
				  deposit.



			 Appropriation of the water, in
				  agreement with the State and potential Protestants.



			 Cleaning of the brackish water i.e.,
				  desalination.






1



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The exploration potential, for the most
		part, will be predicated upon success at the Burgin. There are underground
		access points to explore defined zones and targets adjacent to and accessible
		by the Burgin 1050 level.





2



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3



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TERMS OF
		REFERENCE

(a)
			 The writers of this technical report
				  have been requested by Chief to utilize all the historic data pertaining to the
				  Tintic District and evaluate its importance in the definition of the present
				  assets of the Company. With this definition determined, we are to consider and
				  describe these assets, along with a program to increase their value to the full
				  and optimum extent.


(b)
			 This technical report has been
				  prepared to facilitate the staged financing of the development plans required
				  by (a) above. Corporate decisions made by Chief may require initial private
				  financings followed by a public prospectus. Thus this report has been written
				  to facilitate conversion to N1 43-101-F1 in the eventuality that it will, in
				  whole or part, be utilized for public financing through the auspices of the
				  Canadian Exchange.


(c)
			 The information to be utilized comes
				  from the surviving data from some 140 years of activity within the Tintic
				  District and particularly Chiefs large land holdings. Perhaps more
				  relevantly, the work carried out by previous lessees, namely Kennecott Copper
				  Corporation and Sunshine Mining Company has been referenced. More recently data
				  has been generated by Chief, either singly, or in joint venture with Akiko
				  Resources and Korea Zinc. All these data, primarily drill hole results,
				  metallurgical and process analyses and Water Application and studies can be
				  found in a large filing system, in the offices of Chief, located on the site of
				  the historic mining operations, adjacent to a 800 ton per day process
				  plant.


NOTE:

Throughout this report, we have referred to
		tonnage, resources, reserves and grades that pre-date N1 43  101. In some
		instances, these references are based upon previous workers impressions
		of what has been typical in the past. The inclusion of these data is made to
		identify targets and projects that should be investigated further with current
		equipment and technology. As such, these criteria are understood to be part of
		the Terms of Reference.





4



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PROPERTY
		DESCRIPTION AND LOCATION

Chiefs property in the Tintic Mining
		District comprises some 19,300 acres of patented mining claims and
		approximately 200 acres of non-patented claims. This property is located in
		Juab and Utah Counties of the State of Utah.(See Figure No.2)Of the above claims, approximately 8,500 acres of
		patented ground and 200 acres of non-patented are owned by Tintic Utah Metals
		LLC, which is a subsidiary of Chief. The centre of the property is
		approximately 39° - 57 ́ N latitude and 112° - 20 ́ W
		longitude. The area is dominated by the East Tintic Mountains, a north trending
		fault block range near the eastern margin of the Great Basin.

We have not determined the full extent and
		location of Chiefs claims. For the purpose of this study, there is a
		concentration of effort applied to the East Tintic District and the approximate
		claim boundary there is shown onFigure
		No.3.





5



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6



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ACCESSIBILTY,
		CLIMATE, LOCAL RESOURCES,

INFRASTRUCTURE AND
		PHYSIOGRAPHY

The property area is some 60 miles south
		west of Salt lake City, Utah and is accessible by paved road. Initially south
		on Interstate 15 from Provo to Santaquin, a distance of 15 miles, then west on
		Highway 6 (which bisects the area) a distance of approximately 20 miles. A spur
		from the Denver Rio Grande Western Railway line services the Burgin Mine, which
		is the location of the primary asset development target of the many remaining
		on the property.(See Figure
		No.2)

As previously stated, the topography is
		dominated by north trending ridges with associated valleys. Maximum relief in
		the area is some 10,000 ft with lowest elevations at 5000ft. Precipitation
		amounts to approximately 15 inches per year, equally proportioned between snow
		and rain. Average mid winter temperature (January) is 38°F with a recorded
		minimum of minus 18°F. Mid summer (August) the average temperature is
		74°F and maximum recorded high at 112°F. Vegetation in the claim area
		is typical of semi-arid mountainous terrain. Sage juniper, hedgehog cactus and
		prickly pear dominate on the hotter and drier south facing slopes. Aspen,
		Douglas fir and spruce cover the higher and north facing slopes.

The town of Eureka, altitude 6460ft (1964m),
		population 760, lies adjacent to the claim boundaries. This town has serviced
		previous mining activities. There are a few small towns located eastward on
		Highway 6 through to Santaquin which is located at the junction between Highway
		6 and Interstate 15, and has a population of some 5000.





7



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HISTORY

GENERAL

Ore, in the Tintic District, was initially
		discovered in outcrop in 1869. Within a few years most of the major outcropping
		ore-bodies were being mined and there were 15 -20 exploitation and exploration
		shafts. In addition to the initial Sunbeam lode (small fissure type ore body)
		major ore-bodies (replacement type) were discovered along three main structures
		known as the Gemini (Eureka) the Mammoth and the Godiva-Sioux ore runs (or
		zones). In 1905 the fourth and latest of the ore runs, named the Iron Blossom
		was found by Jesse Knight. Anecdotally, the Mormon, Bishop Knight was led to
		the zone via a vision. However, Knight was familiar with the district and no
		doubt recognized the altered appearance of the surrounding rocks even though
		the ore body did not outcrop. This blind discovery by Knight some
		distance east of the outcropping ore runs, opened up the possibility of further
		deposits to the east. (See sketch below)









8



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By the end of the 1871, three mining camps
		had been established  Eureka, Silver City and Diamond City. The growth
		along the discovered ore runs was steady and most of the early producers were
		still in operation 40 years later. The Mammoth, Martha Washington, Eureka Hill
		and Shavers were amongst the earliest producers. As a result of the high grade
		ores, shipments were sent long distances to smelters in San Francisco, Reno,
		Baltimore and as far as Swansea in Wales.

Even though many claims had been staked in
		the eastern portion of the district (Now East Tintic District) before the turn
		of the century, the only sniff of ore was a small outcrop of
		lead-silver near to the present Eureka  Lily shaft. This outcrop was in
		an area of intense alteration of both sediments and volcanics. All future
		discoveries of major deposits in the East Tintic would be blind ore-bodies,
		based on surface alteration and found by underground geologic interpretation.(See Figure No.3)

E.J. Roddatz became interested in the
		district about 1906 and acquired a major holding in what is now the Tintic
		Standard area. Roddatz reasoned that even though the surface rocks were
		inhospitable, there was a chance of discovery in the Ophir limestone at depth.
		His ideas worked out and, despite the complex faulting and folding in the area,
		and with great perseverance, in 1916, he discovered the Tintic Standard
		deposit. It took two shafts and thousands of feet of drift and winze to make
		the discovery. Roddatz was amply rewarded. The Tintic Standard Mine went on to
		become one of the major lead-silver mines in the world.

The Tintic Standard discovery stimulated
		activity in the eastern part of the district and since the Tintic Standard No.2
		shaft (discovery shaft) was sunk in an area of pyritized volcanics, much of the
		prospecting concentrated on similar alteration zones. The Eureka Bullion, Iron
		King, Copper Leaf, Eureka Lily, Zuma and Apex Standard shafts were sunk in
		similar zones. With the exception of the Eureka Lily, none led to an immediate
		discovery, this would come later after much effort.

Mining geologists attracted by the discovery
		of the Tintic Standard began to study the district. Amongst these was Paul
		Billingsley, who, in the early 1920s, provided the skills that led to the
		framework for modern prospecting. He observed that the volcanic cover in the
		east was pre mineral and altered by various stages of the mineralizing
		solutions contemporaneously with ore deposition. He also recognized that the
		dikes and fissures cutting the volcanics continued at depth into the underlying
		sediments. Based on these ideas a long drive on the 700 level of the Tintic
		Standard Mine was commissioned. The target was projected from a surface
		exposure of a strong alteration zone, along with persistent dikes in the
		volcanics. This exploration work intersected the ore deposit that became the
		North Lily Mine. Similar strategies led to the discovery of the Eureka Standard
		mine.

During World War II, the United States
		recognized that in the event of a long war, new sources of raw material would
		be essential. The US Geological Survey spent 1942  1943 studying the East
		Tintic District. An exploration program seeking blind ore-bodies, was commenced
		towards the end of 1943. The work progressed slowly and by 1945, four targets
		had been defined. One of the targets was the Chief Oxide area. This area was
		centered on a prominent outcrop of oxidized and pyritized volcanics. No major
		discovery was made via the 75ft deep Chief Oxide zone shaft. (It is interesting
		to note that Paul Billingsley, late in the 1920s, had defined the Chief
		Oxide zone as a prime target.) A drift was driven from the Apex Standard and
		exploration holes drilled downwards, This also was unsuccessful. It has been
		subsequently proven that both Billingsley and USGS were correct in their
		thinking. The Burgin deposit lies within the Chief Oxide zone and below the
		Billingsley exploration holes.





9



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In summary, district production slowly
		increased through discovery of new mines and peaked between 1921 and 1930, when
		according to data from the U.S. Bureau of Mines (Morris and Mogensen (1978))
		production for the decade from the combination of the Tintic and East Tintic
		Mining Districts reached 4,250,000 tons. From that peak, production decreased
		to a low of 662,000 tons between 1961 and 1970. Production from the Burgin Mine
		led to a second peak of 1,200,000 tons between 1971 and 1976. Total recovered
		metal from the district between 1869 and 1976 is as follows:



Tons
			 
			 18MMGold
			 
			 2.7MM ouncesSilver
			 
			 270MM ouncesCopper
			 
			 250MM lbs.Lead
			 
			 2.2 billion lbs.Zinc
			 
			 250MM lbs.


Martineau and Potter (1977) report metal
		production from the largest mines in the district and that summary is
		reproduced here in the following table.

METAL PRODUCTION FROM THE LARGEST MINES
		IN THE

TINTIC DISTRICT

TABLE NO.1



MINE
			 
			 TONS(millions)
			 
			 SILVERoz/ton
			 
			 GOLDoz/ton
			 
			 LEAD%
			 
			 COPPER%
			 
			 ZINC%Burgin
			 
			 1.56
			 
			 9.3
			 
			 -
			 
			 10.3
			 
			 -
			 
			 10.1(through 1976)Centennial
			 
			 1.42
			 
			 12.6
			 
			 0.4
			 
			 0.6
			 
			 2.4
			 
			 -EurekaChief
			 
			 3.5
			 
			 15.5
			 
			 0.1
			 
			 6.0
			 
			 0.3
			 
			 2.3ConsolidatedEagle+Bluebell
			 
			 0.67
			 
			 11.7
			 
			 0.1
			 
			 7.5
			 
			 0.2
			 
			 -Iron Blossom
			 
			 0.65
			 
			 26.3
			 
			 0.1
			 
			 9.5
			 
			 0.6
			 
			 -Mammoth
			 
			 1.18
			 
			 11.1
			 
			 0.1
			 
			 1.5
			 
			 1.6
			 
			 -Tintic Standard
			 
			 2.41
			 
			 24.4
			 
			 0.04
			 
			 11.9
			 
			 0.4
			 
			 -






10



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11



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THE TRIXIE
		MINE

The most recent mining efforts have been
		operations from the Trixie Mine.(See
		Figure No.3)The Trixie shaft extends
		to a depth of some 1,300 ft and was active from 1969 to 1992. Operators
		initially were, Kennecott Mining Company, followed for the last nine years by
		Sunshine Mining Company.

During its life the Trixie produced a
		reported 713,478 tons of ore, which was directly shipped (without
		concentration) to the Bingham smelter in Utah for smelter flux. Metals produced
		were 150,048 ounces of gold and 4,670,289 ounces of silver  i.e. grades
		of 0.21 opt Au and 6.5 opt Ag.

This ore was contained in the vertical
		interval from the 1,300 ft level to the 750 ft level.

During 2001, Chief completed a limited drill
		program of six holes drilled from surface into the Jake Gold Zone. The assay
		results from these holes, along with five drilled earlier, were used by a
		Mr. Ray Irwin to estimate a Trixie Mine reserve. It should be noted that
		the available data indicates that the six holes intersected the Jake Zone at
		about 600 ft  700 ft. We have not established the inclinations of these
		holes but are reasonably satisfied they were not vertical. Therefore, the
		estimate should relate to reserves that are above or near the 600 ft
		level.

Mr. Irwin did a series of estimates
		based upon the cutting of high grade intersects. These estimates
		varied from approximately 50,000 tons of 0.55 opt Au, 4.45 opt Ag (cut) to
		50,000 tons of 1.6 opt Au, 4.45 opt Ag.

Also, during 2001, Chief was actively
		engaged in developing the high grade zone (we assume this refers to the Jake).
		Correspondence at the time indicates they were estimating 120,000 tons at + 1.0
		opt Au. We have not yet been able to verify where this reserve was thought to
		be, nor any technical rationale for the estimate. The 600 ft level was to be
		the focus of the proposed mining.

Mining from the Trixie commenced in January
		2002; there had been some development ore produced and stockpiled during 2001.
		On March 28th, 2002 a serious cave-in at the 610 Stope,
		put an end to the mining. Records exist up until the end of February 2002 for
		the mine production. As of this report, we have not been able to find the March
		2002 production information.





12



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Relevant data to the end of February 2002 is
		as follows:



2001
			 
			 Jan. 2002
			 
			 Feb. 2002
			 
			 TotalTons Mined
			 
			 5,384
			 
			 771
			 
			 1,653
			 
			 
			 7,808Tons Milled
			 
			 381
			 
			 3,391
			 
			 4,036
			 
			 
			 7,808Mill Head Grade (Au)
			 
			 0.28
			 
			 0.63
			 
			 0.83
			 
			 
			 0.72Recovery
			 
			 85.9
			 %
			 79.7
			 %
			 80.1
			 %
			 
			 80.1
			 %Recovered ounces
			 
			 91.31
			 
			 1,694.6
			 
			 2,674.5
			 
			 
			 4,460.4Ounces sold
			 
			 
			 
			 
			 
			 
			 
			 
			 3,316.7Ounces in mill circuit
			 
			 
			 
			 
			 
			 
			 
			 
			 1,143.6Total Mining & Milling
				  Cost
			 
			 
			 
			 
			 
			 
			 
			 $
			 694,591Cost per Ton Processed
			 
			 
			 
			 
			 
			 
			 
			 $
			 88.96


In June 2002, Mr. Ray Irwin estimated
		the following reserve as being present above the 750 level i.e. the area of his
		earlier 2002 estimates.



610 Stope
			 18,500 tons
			 
			 
			 0.68 opt Au; 5.05 opt Ag=
			 12,580 ounces Au; 93,500 ounces
				  Ag750 level
			 5,600 tons
			 
			 
			 0.23 opt Au; 0.6 opt Ag=
			 1,290 ounces Au; 3,360 ounces
				  Ag600  610
			 1,750 tons
			 
			 
			 1.85 opt Au; 2.83 opt Ag=
			 3,250 ounces Au; 4,950 ounces
				  AgTOTAL
			 25,850 tons
			 
			 
			 Containing 13,080 ounces Au
				  and101,740 ounces Ag


Without the specific information as to the
		March 2002 production it is difficult to make effective comments. If we assume
		that approximately 4,000 tons were mined and processed in March, then the total
		production tonnage equates to approximately 12,000 tons. Mr. Irwins
		June 2002 estimate brings the total up to 37,850 tons. The discrepancy between
		the original 120,000 tons, Mr. Irwins original 50,000 tons and
		37,850 tons, needs to be field investigated since, in the event of the plant
		becoming operational, this provides an excellent gold target.





13



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BURGIN

After World War II, Newmont Mining Company
		became interested in the Chief Oxide Zone. Operating from the Apex Standard
		No.2 shaft, they carried out exploration seeking blind ore bodies. Their lack
		of success led to a termination of the work in 1948. The last effort they made
		was to drill two churn holes from surface. The first hole (N-1) penetrated a
		thick sequence of carbonate rocks and approximately 50ft of altered dolomite
		containing traces of lead and zinc. Hole number two (N-2) located about 400ft
		northeast of hole number one, intersected nothing of interest and terminated at
		the water table.

The USGS, intrigued with the geology in hole
		number N-1, deepened it to 1600 ft. Fossils of Devonian and Mississippian age
		were identified, which eliminated the previous belief that the rocks were older
		Cambrian strata. This discovery led to the probability of a major north
		trending fault buried under the volcanics. This theory would explain the
		presence of the younger rocks. This fault was considered to be a major thrust
		fault which moved older Cambrian quartzites and shales over the younger
		carbonate rocks.

Bear Creek Mining Co. Ltd., a subsidiary of
		Kennecott Copper, was, at the time, active in the district. Their interest
		extended to the Chief Oxide Zone. Their initial work was managed by
		Mr. William Burgin. Mr. Burgin died in an airplane crash in 1955 and
		the eventual Chief Oxide Zone discovery was named in his honour. In 1957 what
		is now the Burgin No.1 shaft was sunk some 50ft east of the Chief Oxide shaft
		and surface drilling commenced. Centennial Development Company (the Contractor)
		collared the shaft on January 30th 1957 and had reached the water
		table some 1110ft below on July 20th 1957. The 1050 level was
		established approximately 20ft above the water level and an exploratory
		cross-cut was sent off in a westerly direction. The cross-cut had two
		objectives.


			 To enter the postulated East Tintic
				  thrust fault



			 To investigate Bear Creek surface
				  hole ET1 which was an extension of an old USGS hole that had intersected
				  lead-zinc mineralization


Both objectives were achieved. The thrust
		fault was intersected some 1300ft west of the shaft. Drifts were expanded into
		and down the fault and intersected numerous zones of lead-zinc oxide and lead
		carbonate. For the next three years drilling above and below the 1050 level
		continued to increase the lead-zinc silver deposit that became the Burgin Mine.
		A production decision was made in 1963. The plan called for 75 tons per day of
		direct shipping ore initially, rising to full production of 500 tons per day by
		1965. The basis of the positive decision was 1,330,000 tons of lead - silver
		 zinc ore blocked out by drilling from the 1050 level. As a matter of
		some significance, at the time of the decision, some water flow during mining
		was anticipated, initially 3000 gpm  4000 gpm, decreasing to 2000 gpm
		 3000 gpm. during production.





14



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As initially planned, the Burgin Mine would
		consist of a new (Burgin No.2) shaft which would have production levels at
		1050ft, 1200ft and 1300ft. A pumping station and associated water storage would
		be installed at 1350ft and the shaft would be terminated at the 1500 ft level.
		Rehabilitation of the near-by Apex No2 shaft to the 1000ft level and two
		connecting drifts would be provided for escape and ventilation purposes. Raise
		bored holes, winzes and raises would complete the main mine facilities.

This plan would be drastically amended, with
		negative impact, due to the problem that was to plague operations at the Burgin
		Mine.

One other feature of the initial operational
		plan was the elimination of a concentrating process plant. The preponderance of
		oxide ore presented metallurgical challenges that did not justify the risk of
		the additional cost. Production would come from high grade silica-laden silver
		lead-zinc ore that would be directly shipped to the Utah Kennecott
		smelter.

During 1965, with the Burgin No.2 shaft at
		1330 ft, an inflow of 140°F water at 1400 gpm was encountered in the
		southeast corner of the shaft. Despite intensive dewatering attempts, the flow
		was not stopped. As a result, a decision was made that would negatively impact
		all future operations. Production would commence from the 1200 ft level and
		wells would be sunk from the same level to dewater the workings. Prior to this
		all dewatering had been through a single, deep well north of the shaft. The
		installation of wells on the 1200 level met only limited success and the
		ore-body was not adequately dewatered. It wasnt until a second surface
		well came on-line that the shaft inflow was reduced, and the workings became
		moderately drier.

Production commenced from the 1200 level and
		this hampered the ongoing deeper development. A decision was made to excavate
		the 1300 level station at the 1260 level. This decision also had a serious
		impact on future operations since it decreased by 40% the ore available for
		above-level extraction. During 1966, based upon test-work by Kennecott, a 500
		ton per day flotation concentrator was authorized and installed at the Burgin
		Mine. This was partially built prior to the 1967 strike and completed shortly
		after the strike was over.(See Figure
		No.4)

In 1967, Kennecott was informed by the U.S.
		Steelworkers of America, that they intended to negotiate one contract for all
		workers and the Burgin Mine, under Bear Creek Mining, was to be part of this
		negotiation. By mid July, negotiations had not been successful and a general
		strike commenced. This strike continued until April of 1968 and had a huge
		influence on all subsequent operations. During the strike, the water inflow had
		increased to 4400 gpm. A small sump on 1200 level pumped water to the much
		larger 1050 level sumps, (which had replaced the planned 1350 pumping level).
		From the 1050 level, the water was pumped to the surface for disposal. The few
		supervising staff available to work during the strike were simply overwhelmed
		with the tasks needed to maintain and improve mineability. As a result, the
		mine conditions degenerated considerably, creating wet, unstable ground in all
		of the existing and immediate future stoping areas. Working areas collapsed. A
		back-log of sand filling, immediately prior to





15



---



the strike, created stress on a lot of the
		more important pillars with resulting damage to adjacent working stopes and
		future pillar recovery efforts.

After the strike, it was obvious that the
		dewatering plan was not working. Stope development was delayed; stopes were wet
		and the ground heavy. A lot of the broken ore was removed by eliminating the
		fill method and applying block caving techniques. This resulted in cave
		extensions to the shaft which created shaft movement. Given below is a year by
		year summary, along with the more important lessons that will be of huge
		benefit to any future mining at the Burgin.

1970:Concentrator increased to 800 tpd but it was impossible
		to produce sufficient ore due to a shortage of stopes. There was increasing
		pressure from the top level management to attain the 800 tpd, 168,000 tons
		mined (655 tpd).

1971:140,000 tons mined at 11.8% Pb, 9.4% Zn and 9.0 oz/ton
		Ag. (545 tpd) Most of the ore came from newly discovered zones. Problems
		created by the difficult mandate to feed the concentrator were a lack of grade
		control and implementation of orderly well conceived mine plan.

1972:203,000 tons mined (815 tpd); the increase was due to
		more efficient dewatering wells. During the year the 1300 level was brought on
		stream. Dewatering methods were now beginning to work, an average of 8,850 gpm
		were pumped. The dewatering plan consisted in the main, of deep wells and
		advance drain holes from development headings.

1973:197,000 tons mined (785 tpd) but, as in previous years,
		the bad ground (water) and hot water flows limited access to the high grade
		areas and low grade ore was mined to maintain the budgeted production rates.
		Excessive heat and water inflows occurred at the west end of 1200 level, near
		the intersection of the Eureka Standard and East Tintic thrust fault. All
		headings here were stopped at the contact with the sanded dolomite and shales
		and the footwall of the thrust fault, whilst additional pumping facilities were
		put into place. It was also noted at the time that the quartzite hanging wall
		was adequately dewatered. Therefore it was decided to abandon the footwall
		headings and drive around them in the hanging wall. It was obvious from this
		work that the water flows were at the footwall contact. The quartzite deep
		within the hanging wall had been effectively dewatered.

1974:162,000 tons (650 tpd) of higher grade with excellent
		profit.

1975:Production started to drop due to decreasing ore above
		the 1300 level i.e. more production was required from below the shaft stations
		with inherent problems.

1976:Decrease in production due to lack of reserves. The
		neighboring Trixie Mine started to produce high grade direct shipping ore,
		which alleviated the push for production.





16



---



1977:Although the Burgin Mine still carried in excess of one
		million tons below the 1300 level a decision was made to suspend operations.
		This was based upon the expected costs for a completely new shaft and power
		system, and the expectations that the recently re-opened Trixie Mine would take
		over the Division production duties.





17



---





LESSONS LEARNED AT BURGIN

(1)
			 Despite several profitable years,
				  the overall result was a loss. With hind sight, this loss could have been
				  avoided. Due to pressure from the Kennecott head office and the property owners
				  (Chief) production was taken from areas that had not been dewatered properly
				  nor had effective plans been implemented to minimize dilution (grade control).
				  Additionally, in this respect, it should be noted that Kennecott considered
				  Bear Creek to be a separate Division and all costs incurred in the entire
				  district were carried by that company. These extraneous costs created the loss
				  situation, not necessarily the Burgin Mine operation.


(2)
			 Excess drift-work, including roof
				  support and expensive heading dewatering, had been incurred due to the common
				  practice of footwall development. In this case the use of the hanging wall for
				  all major drifts and development would have been much better.


(3)
			 The mine management felt that the
				  initial concept of a 1365 drainage level and drainage holes, ahead of the
				  workings, was an excellent one. Later, experience demonstrated that a
				  combination of drainage holes and deep wells was very effective. The Kennecott
				  decision to start mining when the shaft






18



---



was stalled at the 1330 level  due to
		water, was the start of all the subsequent problems at the Burgin Mine.

(4)
			 A thorough knowledge of the
				  hydrology is essential and an effective plan designed and allowed to be
				  implemented, would have made a tremendous difference in the profitability of
				  this venture.


(5)
			 All the problems from the end of the
				  1968 strike to closure in 1978 prevented an optimum result. The problems, all
				  due to water, were caving with consequential dilution, ore loss and heavy
				  ground in work places. Water saturated sanded dolomites along the contact of
				  the East Tintic fault and large areas of the main ore zone that had not been
				  adequately dewatered were a persistent concern. Experience during the life of
				  the Burgin had shown that heavy ground and sanded dolomites could be
				  effectively worked once the rock was dewatered.






19



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GEOLOGICAL
		SETTING

The East Tintic Range is a north trending
		fault block range near the eastern border of the Basin - range province. Its
		shape and orientation is controlled by northerly trending Basin-range normal
		faults, formed late in the geological history of the range. The mountains are
		underlain by Precambrian and Paleozoic strata that underwent complex folding
		and faulting during the Sevier orogeny of Cretacious age. In Oligocene and
		Miocene time the deformed strata were intruded by stocks and volcanic rocks
		buried the youthful land forms.

The Precambrian rocks exposed in the region
		are predominantly quartzite, argillite and brown dolomite belonging to the Big
		Cottonwood Formation. The Upper Proterozoic rocks total approximately 1650 ft
		in thickness.

The Paleozoic rocks which range in age from
		Cambrian to Mississippian unconformably overlie the Precambrian and are
		estimated to be approximately 12,000 ft thick. Morris and Mogensen comment,
		that in general the Paleozoic rocks are about 60 percent carbonate rocks, 30
		percent quartzite and 10 percent shale. The base of the Paleozoic section is
		the Cambrian age Tintic Quartzite. This, more or less, 3,000 ft thick unit is
		mostly buff quartzite but consists of conglomerate near the base and phyllite
		beds near the top. The Tintic is not normally known to host replacement
		ore-bodies, but is host to a number of the narrow, precious metal rich, fissure
		vein deposits.

Overlying the Tintic Quartzite is the Ophir
		Formation of limey shales and limestones of the Middle Cambrian to Upper
		Mississippian. The Ophir is an important ore host in the East Tintic district.
		Above the Ophir Formation, sandstones and shale are distinctly subordinate to
		carbonate rocks. All of the units above the Ophir, host at least small
		replacement deposits, but most of the mineral production came from five units
		as follows:


			 The 370 ft thick Ophir Formation
				  ($200 million).



			 450 ft thick Bluebell Dolomite ($135
				  million).



			 510 ft thick Ajax Dolomite ($60
				  million).



			 900 ft thick Deseret Formation ($47
				  million).



			 Approximately 2,700 ft thick Tintic
				  Quartzite ($35 million).


The strong mature relief, developed on the
		eroded sedimentary strata in post Mississippian time, was buried during the
		Oligocene, by ash comprising the Packard Quartz Latite which has been
		correlated with the quartz monzonite Swansea stock and associated dikes.

The Packard Quartz Latite is overlain by a
		sequence of latite flows, tuffs and agglomerates which make up the Tintic
		Mountain Volcanic Group. These rocks formed a large composite volcanic cone
		that was intruded by many stocks, plugs and sills of latite and monzonite
		porphyry of the Sunrise Peak Stock.





20



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The youngest Oligocene eruptive rocks in the
		district belong to the Laguna Springs Volcanic Group. The volume of Laguna
		Springs tuffs and flows is small compared to earlier volcanic events. The
		termination of their eruption was marked by the intrusion of the Silver City
		stock, a monzonite porphyry. Associated with the Silver City stock are numerous
		monzonite porphyry plugs and dikes. Locally, many pebble dikes, which served as
		feeders for ore forming hydrothermal fluids, are associated with the intrusives
		of the Silver City stock.

During and after intrusion of the Silver
		City stock large volumes of hydrothermal fluids coursed through the faults and
		breccia zones of the region and produced extensive regions of hydrothermal
		alteration. These solutions changed in chemical character over time and
		ultimately caused dolomitization, propylitization, argillization,
		silicification, calcitization and sericitization of country rocks. During later
		stages of fluid evolution they deposited the ores for which the district is
		famous.

During the Sevier orogeny, three or more
		superimposed thrust faults and many associated high angle faults were produced
		in the sedimentary pile. None of the thrust faults outcrop within the district,
		but can be traced over long distances in adjacent regions. Paleozoic rocks were
		displaced as much as 100 miles eastward over younger Paleozoic and Mesozoic
		strata. During later stages of movement the strata and thrusts were crumpled
		into asymmetric anticlines and synclines with amplitudes of 3  4.5 miles.
		During the folding, small thrust faults developed on the limbs and crests of
		the structures.

High angle faults, associated with the
		folding include a conjugate system of northeast and northwest trending wrench
		faults showing mainly horizontal displacement.

A series of late faults cut the intrusives,
		volcanics and sediments and can be traced northeasterly across the district.
		Locally, the fault planes are occupied by pebble dikes. These late faults show
		little displacement, typically on the order of a few tens of feet and in most
		instances dip steeply to the west. These faults according to Morris and
		Mogensen (1978) were the primary conduits for ore-depositing solutions and
		where they localize ore shoots the wall rocks exhibit concentric zones of
		hydrothermal alteration.





21



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DEPOSIT
		TYPES

The mined ore-bodies occur as; replacement
		deposits, replacement veins and fissure veins, as well as stockworks and
		disseminated deposits. Replacement deposits have been the source for more than
		90% of the ore produced in the district. They occur predominately, in carbonate
		rocks and may be described as pods, pipes, columns and irregular masses. They
		range from small masses of less than a ton to masses containing more than
		2,000,000 tons. They are irregular in form and commonly connected one to
		another by narrow stringers or veins of ore. Some deposits show a clear
		localization by structure, whilst others are seemingly unrelated to
		structure.

Replacement deposits in the main Tintic
		district occur in five linear zones known as ore runs. Over much of the strike
		length of the ore run it could be expected to be more or less rod shaped, less
		than 100 feet wide and much less than 100 feet high. Much if not most of the
		volume of the ore run would consist of un-replaced wall rock. These are
		geometries that are very difficult to quantify with exploration. Most of the
		mines carried very small tonnages of proven reserves and relied on continuous
		underground exploration. Much of this work was completed by lessors.

The ore-bodies in the East Tintic District,
		including the Burgin, show much more structural control. The major replacement
		deposits are irregular masses localized by the intersections of northeast
		trending fissure zones and small thrust faults with porous, permeable and
		reactive units.

Replacement veins were mined most commonly
		in the pyrometasomatic rocks near the margin of the Silver City Stock, where
		ore replaced the fault gouge and breccia and locally replaced reactive beds.
		Ore shoots were typically mined over a strike length of 300 feet or so,
		although the Dragon vein was mined for more than 2,500 feet. They are sometimes
		pipe like in form and up to 45 or 60 feet in width. Where they are tabular they
		commonly have dimensions of 300 feet vertical and horizontal and 4.5 to 6 feet
		wide. These tabular deposits were far fewer in number than fissure veins, but
		jointly produced a similar amount of ore as the fissure veins, (approximately
		5% each).

Fissure veins occur in a myriad of short
		faults typically cutting massive siliceous rock units including quartzite,
		quartz monzonite, monzonite porphyry, latite, silicified tuff and agglomerate.
		The ore shoots are commonly less than about three feet wide and several hundred
		feet long. Most of these deposits are less than 300 feet long (Morris and
		Mogensen, 1978). These mines typically produced less than 400,000 tons of ore
		from a number of much smaller discrete ore shoots.

The ores in the district consist largely of
		galena and sphalerite with variable amounts of acanthite, argentite,
		tetrahedrite-tennantite, enargite-famatinite, proustite, hessite, calaverite,
		native gold and silver and a wide variety of uncommon copper, lead, silver and
		bismuth sulfosalt minerals. The above mentioned minerals occur in a gangue that
		is commonly siliceous, ranging in texture from coarsely crystalline quartz near
		the Silver City stock to fine grained jasperoid even flinty jasperoid in the
		far north and distal parts





22



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of the district. The gangue also includes
		abundant barite, calcite, dolomite and rhodochrosite. Deep oxidation above the
		water table reaches depths of 900-2,200 feet Morris and Mogensen (1978) in the
		sediments and has produced a variety of sulfate, carbonate, silicates,
		arsenates, antimonates, manganates and other mineral varieties.

The ores intergrade from one type to another
		and not uncommonly, do so within a single mine. Morris and Mogensen (`1978)
		provide the following classification of ore types:

1.
			 Lead ores containing 5-50 percent
				  lead and as much as 50 ounces of silver per ton.


2.
			 Siliceous lead ores containing a few
				  percent lead with minor zinc and copper in a gangue of more than 70 percent
				  silica. Silver values are variable but locally high grade.


3.
			 Siliceous silver ores containing at
				  least 20 ounces of silver per ton and less than 0.15 ounce of gold per
				  ton.


4.
			 Lead-zinc ores containing 5-15
				  percent of each lead and zinc and 8 10 ounces of silver per ton.


5.
			 Copper-gold ores containing a few
				  percent or more of copper, 10 - 20 ounces of silver per ton and commonly more
				  than 0.5 ounce of gold per ton.


6.
			 Gold telluride ores containing minor
				  tetrahedrite and enargite but as much as 2,500 ounces of gold per ton.


7.
			 Siliceous lead-copper ores
				  containing a few percent of copper and lead and occur where copper ores give
				  way to lead ores.


Mineral zoning in the district is
		pronounced. Deposits in the south were most valuable for copper and gold.
		Farther to the north, lead and silver replace copper and gold in importance
		and, are themselves, subsequently replaced in importance by zinc in the
		northern-most deposits. At the Ball Park deposit to the north of the Burgin
		Mine, silver grades are most commonly less than an ounce and zinc is typically
		more abundant than lead. In the East Tintic District, metal zonation similar to
		that seen from south to north across the district can be seen from the deeper
		parts of some deposits to the shallower regions. The North Lily Mine produced
		gold and copper ores at depth and moderate grade lead and silver ores from the
		central and northern parts of the mine. Morris and Mogensen (1978) report the
		Tintic Standard Mine produced gold-copper ore at depth, silver-lead ore from
		the central part of the mine and zincian manganese ore from the upper
		northeastern levels of the mine. The Burgin Mine ores contain lead, zinc, only
		moderate quantities of silver and insignificant amounts of copper and gold.
		Rhodochrosite is abundant in Burgin ores in contrast to its relative rarity in
		other East Tintic ore-bodies.





23



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THE NEW BURGIN
		MINE

BACKGROUND

The Burgin Extension was initially explored
		by Bear Creek Mining Corp., (Kennecott) during the period they mined the Burgin
		down to the 1300 level i.e., 1966  1978. The north-south trending East
		Tintic thrust fault is present all along the east margin of the district.
		Movement along this structure has placed Cambrian quartzites over Mississippian
		and Devonian carbonate rocks. Where northeast trending structures intersect the
		base of the thrust plate, mineralization has been emplaced in zones of
		brecciation and also as a direct replacement of favourable stratigraphic
		horizons. The Burgin ore bodies and ore lenses are located in or near the
		thrust fault intersection with three known northeast trending tear faults. The
		tear faults have provided the open conduits for ore bearing solutions to rise
		and invade the broken ground in the thrust zones. The receptive stratigraphic
		units are the limestones and dolomites in contact with the thrust fault and the
		brecciated quartzites. In the vicinity of the ore and close to faults, the
		limestone beds have been dolomitized and commonly contain sanded
		caverns.

The Burgin Extension ore-body is irregular
		in size, shape and orientation. In plan view it is roughly parabolic and it
		plunges to the west at approximately 35°. Included in the ore-body is an
		isolated pod located in the footwall of the thrust fault below the 1300 level
		(4,395 el.) known as the Pothole ore-body.(See Figure No.5)

The primary metallic minerals include
		argentiferous galena, sphalerite, pyrite and some silver sulfosalts. Ore grades
		vary but in some areas are as high as 70% combined Pb/Zn, indicating almost
		complete replacement by sulfide minerals. Lower grade mineralization in general
		extends beyond the ore/waste boundaries, in essence forming a halo around the
		main ore-body. Gangue minerals include barite, rhodochrosite, calcite,
		dolomite, quartz and a number of clay minerals. There does not appear to be any
		oxidation below the water table.





24



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25



---



Since the initial drilling, the Extension
		has been subjected to periods of intermittent exploration. These campaigns have
		been carried out by reputable mining and development companies. The sequence of
		development is shown in the following table, along with reserve/resource
		estimates made.

HISTORICAL RESOURCE/RESERVE
		ESTIMATES

TABLE NO.2



Year
			 
			 Company
			 
			 Resource(Reserve)
			 
			 AgOpt
			 
			 Pb%
			 
			 Zn%
			 
			 CommentsJan 1978
			 
			 Bear CreekJan 1981
			 
			 Sunshine Mining
			 
			 1,623,000
			 
			 12.35
			 
			 16.26
			 
			 7.12
			 
			 Total Resource no
				  dilutionJan 1982
			 
			 
			 
			 1,741,000
			 
			 12.80
			 
			 16.14
			 
			 6.85
			 
			 Total Resource zero
				  dilutionJan 1984
			 
			 
			 
			 1,624,000
			 
			 17.46
			 
			 18.88
			 
			 5.31Dec 1985
			 
			 
			 
			 1,150,000
			 
			 22.20
			 
			 18.78
			 
			 6.83Jan 1988
			 
			 
			 
			 1,284,000
			 
			 22.57
			 
			 19.54
			 
			 5.53Dec 1989
			 
			 Pincock, AllenAnd Holt
			 
			 1,032,000
			 
			 23.10
			 
			 26.65
			 
			 8.74
			 
			 Based on 47 holes a
				  netoperating revenue of
				  $65/tonJuly 1996
			 
			 D.Tschabrun P.E.For Akiko Gold
			 
			 1,339,000
			 
			 20.81
			 
			 15.72
			 
			 6.49
			 
			 Based on 81 holes and
				  a$25.22/ton cut-offNov 1997
			 
			 D. Tschabrun P.E.For Chief
			 
			 1,406,000
			 
			 16.27
			 
			 12.76
			 
			 5.72
			 
			 Based on 105 holes and
				  a$25.22 cut-off


There have been several feasibility
		studies carried out on the Extension. The earliest was completed by SMC
		in 1988, and each subsequent one has been an update with a slight change in
		resource input as indicated in the evolution table above.(Table
		No.2)In every case, the fundamental
		concept of separate saleable lead and zinc concentrates has been assumed. Our
		conclusions relative to this are outlined in the Metallurgical section below,
		but in summary, we feel that it is not a reasonable assumption. There does
		however appear to be an area of the deposit where the Pb/Zn ratio is conducive
		to the production separate concentrates. This area will be further defined and
		investigated in the work program suggested in this report.(See Figure No.5)

The most recent feasibility update was
		completed by Mine Development Associates (MDA) of Reno, Nevada in 2001. They
		applied the two concentrate concept to the Tschabrun resource model, and
		assumed mine water would be pumped to adjacent old workings.

MDAs main parameters are shown below
		for historic purposes. We are of the opinion that Chief should, rather than a
		simple new update, pursue an open line of thinking leading to a new Feasibility
		Study for the Burgin. The current Resources, which were considered by MDA to
		produce a positive cash flow, need augmentation as to Resource





26



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criteria. A drill program is recommended
		that also potentially increases the overall Resource.



MDA STUDY
			 Capital Cost
			 $44,819,000(2001)
			 Operating Cost
			 $67.44 per tonDaily Production
			 800 tons per day


FUTURE
		PROGRAM

We are recommending that an underground
		drill program be conducted from the Burgin 1050 level, accessed via the Apex
		No. 2 shaft. Surface drilling has been discounted due to the inability to
		ensure the drill penetrates the desired ore zone and exactly twins existing
		holes. Prior to commencement of this program it will be necessary to provide
		electrical power to the mine and clean up the surface facilities. The shaft and
		1050 level will need inspection and rehabilitation as necessary. Ventilation,
		communications and safety facilities will all require attention.(See Figure No.5)

This drill program will provide technical
		data to complete those essential pre-requisites to a Bankable Feasibility
		Study.

1.
			 Twin old holes to provide drill
				  campaign comparisons and statistical analysis and to facilitate a N1 43-101
				  Resource estimate.


2.
			 Metallurgical samples to define
				  concentrate characteristics and the different metallurgy between the indicated
				  Lead and Zinc dominant zones.


3.
			 Geotechnical and hydrological data
				  for mine planning and dewatering.



			 Logging of ore-body structure



			 Packer testing of rock
				  permeability



			 Installation of maintaining
				  wells


This work will be the initial investigation
		leading to an appropriate groundwater mitigation plan.

Drill core will be split to provide high
		zinc metallurgical samples from the ore adjacent to and parallel to holes B-122
		and CB-9. High lead metallurgical samples are to be acquired by drilling
		adjacent to and parallel to holes SB-14 and B-68. A third site proposed for
		acquiring high lead metallurgical samples is by drilling to the southeast from
		the site of hole CB-30. The first four holes mentioned above will also provide
		twins of earlier drill holes that can be used to improve the confidence level
		of the resource estimates based on the earlier drilling.

Examination of drill plans and sections,
		leads to the conclusion, that in several locations, the drill hole pierce
		points of mineralization are somewhat less than ideally spaced or that the
		drilling was stopped short of the mineralization. It is therefore proposed that
		additional drill holes be located to close some of the larger gaps in the drill
		pattern.





27



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The review of earlier drilling identified
		several deeper holes in the western part of the drill pattern that appear to
		have intersected a relatively untested deeper zone of mineralization located
		approximately 250 to 350 below the mineralization explored in
		detail by the earlier drill programs. The intersections are shown in the table
		below.

DEEPER DRILL HOLE INTERSECTIONS, WESTERN
		SIDE OF BURGIN DRILL PATTERN

TABLE NO.3



Hole Identifier
			 
			 Width (ft)
			 
			 Silver (g/t)
			 
			 Lead (%)
			 
			 Zinc (%)SB-1
			 
			 73.0
			 
			 20.57
			 
			 18.7
			 
			 6.8SB-6
			 
			 26.7
			 
			 25.14
			 
			 15.4
			 
			 5.28SB-9
			 
			 12.0
			 
			 11.07
			 
			 17.1
			 
			 0.18


A number of drill holes in the recommended
		program have lengths extended to provide a check for additional mineralization
		at the potentially productive horizon/structure intersected.

The drill holes recommended are listed in
		the table below. It is suggested that HQ gauge core be used to maximize the
		amount of sample material available.

RECOMMENDED DRILL HOLES

TABLE NO.4



Identifier
			 
			 Northing
			 
			 Easting
			 
			 Direction
			 
			 Inclination
			 
			 Length (ft)1
			 
			 31238
			 
			 25898
			 
			 2850
			 
			 -700
			 **
			 5002
			 
			 30880
			 
			 25815
			 
			 00
			 
			 -900
			 
			 5003
			 
			 30929
			 
			 25910
			 
			 800
			 
			 -380
			 
			 3504
			 
			 31028
			 
			 25693
			 
			 1300
			 
			 -600
			 
			 650
			 (400)*5
			 
			 31028
			 
			 25693
			 
			 950
			 
			 -650
			 
			 650
			 (400)*6
			 
			 31000
			 
			 25988
			 
			 00
			 
			 -900
			 
			 1507
			 
			 31229
			 
			 25906
			 
			 00
			 
			 -900
			 
			 650
			 (400)*8
			 
			 31025
			 
			 25693
			 
			 1870
			 
			 -610
			 
			 700
			 (525)*9
			 
			 31359
			 
			 25712
			 
			 00
			 
			 -900
			 
			 600
			 (400)*10
			 
			 30938
			 
			 25903
			 
			 00
			 
			 -900
			 
			 65011
			 
			 31359
			 
			 25712
			 
			 900
			 
			 -650
			 
			 600
			 (500)*Sub-Total
			 
			 
			 
			 
			 
			 
			 
			 
			 
			 6,000Footage Contingency
			 
			 
			 
			 
			 
			 
			 
			 
			 
			 1,200Total
			 
			 
			 
			 
			 
			 
			 
			 
			 
			 7,200


*
			 Length required if not extended to
				  check for deeper mineralization


**
			 Implies an inclination to penetrate
				  rock materials below the level of the drill hole collar.


Following the drilling, we anticipate
		potential further project financing using public methods. This will require a
		N1 43-101 Resource and subsequent Scoping Study completion. The last three
		months of a nine month program is allocated for these studies at a cost of
		$100,000.





28



---



The estimated budget for the Burgin program
		is as follows:



Power to Mine
			 
			 
			 $
			 10,000Surface clean-up
			 
			 
			 
			 3,000Shaft and Mine
				  Inspection
			 
			 
			 
			 10,000Rehabilitation (unknown 
				  pending inspection)
			 
			 
			 
			 Drilling
			 7,200 ft @ $35 per ft
			 =
			 
			 $
			 252,000Assays
			 2700 @ $24 per assay
			 =
			 
			 
			 65,000Supervision and Geology
				  (Miscellaneous expenses)
			 =
			 
			 
			 100,000Hydrology
			 =
			 
			 
			 50,000Resource and Scoping
				  Study
			 =
			 
			 
			 100,000Sub-Total
			 =
			 
			 $
			 590,000Contingency @ 20%
			 
			 
			 $
			 120,000BUDGET
			 
			 $
			 710,000


MINERAL
		PROCESSING

The Burgin Mine and Concentrator was
		operated by Bear Creek Mining Co., a subsidiary of Kennecott Copper
		Corporation, from 1963 to July, 1978. They produced separate lead and zinc
		concentrates using differential flotation from both sulphide and oxide ores.
		Ore was processed at rates which varied from 400 tpd to 900 tpd depending upon
		grindability of the mineralization and availability of the mine and plant. A
		conventional crushing circuit preceded two ball mills in parallel which fed a
		lead flotation circuit followed by a zinc flotation circuit, concentrate
		thickening, filtering, concentrate load-out, and tailing disposal(See Figure No.6). The second ball mill was added in June, 1975 in order
		to raise milling capacity from 400/500 tpd to 700/900 tpd. In the final years
		of operation, predominantly sulphide ores were processed.





29



---











30



---



During its tenure, Kennecott processed ore
		grading minor Au, 12.7 opt Ag,, 0.03% Cu, 13.9% Pb, and 15.5% Zn. Overall
		recoveries were reported as 83% for silver, 81% for lead, and 68% for zinc.
		Kennecott also conducted several investigations into reasons for poor
		metallurgical performance with respect to recovery of zinc and the grade of the
		zinc concentrate, as well as losses of lead and silver. Consistent conclusions
		as a result of mineralogical examination of samples of concentrates and
		tailings were:


			 The intimate association, by
				  attachment, of iron-poor sphalerite with pyrite and non-sulphide minerals
				  (rhodochrosite, quartz, barite, and dolomite), either as inclusions or as
				  surface middlings;



			 Rimming of sphalerite by fine
				  galena;



			 Small inclusions of galena in
				  sphalerite;



			 Coating of sphalerite by dickite
				  (clay).


There are, in the literature (USGS
		Professional paper 614A by Radke, et alia.) references to two types of galena
		in the Burgin Mine: Type 1, coarse-grained and silver-rich, and Type 2,
		fine-grained and silver-poor. Such occurrence is not unusual in polymetallic
		mineralization.

Although a regrind circuit was installed in
		the Concentrator, no recorded evidence of its use has been found, either in the
		lead circuit or the zinc circuit, though it must have been used at some point.
		There is ample comment in Kennecotts observations about the necessity for
		finer liberation, particularly for sphalerite, but no assessment of an economic
		grind to flotation has been found. Consequently, it appears that the
		Concentrator was operated on the basis of achieving maximum hourly feed rate
		through the ball mills, consistent with the output of saleable concentrates,
		but at the expense of optimization of the flow-sheet and reagent use.

As part of a Scoping Study, capital costs
		should be developed for re-habilitation, repair, and re-commissioning of the
		Concentrator and ancillary facilities. The Concentrator itself is dry and
		relatively clean and the remaining installed equipment appears to be in good
		condition, though in some areas it is incomplete. In contrast, any equipment
		which is located outside, will require major work. The property has suffered
		from theft, particularly in electrical cabling, piping, tools, and small
		equipment, e.g., compressors and welding machines. During the recent Trixie
		gold ore milling it was not possible to process more than 240 tons per day, due
		to limitations in the classification circuit. The gravity circuit consisting of
		a Knelson concentrator that recovered free gold, has been sold; its replacement
		cost would be approximately $120,000, failure to provide a gravity circuit for
		similar ores would reduce gold recovery by about one-third. Some equipment has
		been sold to pay creditors. A detailed inventory and assessment of deficiencies
		has not been undertaken within the scope of this report. Every part of the
		plant will have to be cleaned and made safe (i.e., necessary platforms, stairs,
		cranes, lighting, and heating, etc.)





31



---



prior to any new work. It is assumed that
		existing process equipment will not have to be replaced, except as noted.

In summary:


			 All conveyors, crushers, screens,
				  feeders, storage silo, and dust collector should be inspected for structural
				  integrity, where applicable, and mechanical operability, followed by repair and
				  replacement of missing components; all motors should be inspected and
				  tested;



			 The ball mill inside the plant, its
				  motor/drive components, and its mill lining should be inspected by the
				  OEM/Vendor and re-commissioned;



			 The ball mill outside the plant, its
				  motor/drive components, its mill lining, and its foundation will require major
				  work, again by the assigned millwrights, and re-commissioned;



			 The regrind mill, its motor/drive
				  components, and its mill lining should be inspected by the assigned millwrights
				  and re-commissioned;



			 The inventory of pumps and cyclones
				  in the grinding circuits will have to be completed and tested, and
				  piping/launders replaced or added to where necessary;



			 The single bank of flotation cells
				  (the only one remaining from later operation on Trixie ore) will have to be
				  assessed for the required residence time to produce a bulk concentrate, and
				  most likely added to, plus the purchase of some additional cells for
				  cleaning;



			 Both thickener rakes and supports
				  will have to be assessed for structural integrity and in one case re-built;
				  mechanisms will have to be tested and re-commissioned;



			 Both concentrate filters will have
				  to be re-built using the remaining useable components with additional bags;
				  vacuum pumps will have to be tested;



			 The inventory of process pumps will
				  have to be completed and tested, and piping/ launders replaced or added to
				  where necessary;



			 All small motors and drives will
				  have to be inspected, tested, and added to where necessary;



			 The concentrate load-out facility
				  will have to repaired (concrete) and rail access restored;



			 The supply of fresh and process
				  water will have to be restored and the reservoir cleaned out;






32



---




			 The pipeline for disposal of
				  tailings will have to be surveyed and most likely replaced; the lining for the
				  pond will also require repair;



			 The system for supply and
				  distribution of power from the Public Utility, including the motor control
				  centres and control room, will have to be inspected, repaired as necessary, and
				  certified;



			 The building which housed the
				  metallurgical/assay laboratories will have to be inspected, repaired, and
				  re-equipped for on-site test-work and assaying, including environmental
				  protection and monitoring;



			 The inventory of tools, small
				  equipment, and vehicles for mechanics and electricians will have to be
				  replaced;



			 The offices will have to be cleaned
				  out and made habitable with plumbing, toilets, heat, and light.


An order-of-magnitude estimate of cost for
		re-habilitation, repair, and re-commissioning of the Concentrator and ancillary
		facilities to produce a bulk flotation concentrate at the rate of 800 tons per
		day is, in Year 2005 constant dollars, US$ 3,500,000 to US$ 5,000,000.

METALLURGICAL
		TESTWORK

During this investigation, a report by
		Dawson Metallurgical Laboratories, Inc. in 1997 and commissioned by Chief
		Consolidated Mining Company was sourced in which some differential flotation
		test-work at batch-scale showed that a separate concentrate for lead and
		silver, with attendant gold in minor quantities, and good recoveries could be
		produced, followed by a zinc concentrate of dubious grade with poor recoveries.
		The origin of the ore samples was unclear and the head grade was higher than
		for Kennecotts Life-of-Mine average:

FLOTATION TESTWORK

(Pb/Zn Ratio 6:1)

TABLE NO.5



Assay
			 
			 Recovery, %Product
			 
			 Wt %
			 
			 Ag,oz/ton
			 
			 Pb,%
			 
			 Zn,%
			 
			 Ag
			 
			 Pb
			 
			 ZnPb Concentrate
			 
			 31.45
			 
			 56.17
			 
			 62.99
			 
			 3.98
			 
			 85.3
			 
			 90.1
			 
			 35.5Zn Concentrate
			 
			 4.50
			 
			 16.00
			 
			 32.79
			 
			 39.89
			 
			 3.5
			 
			 6.7
			 
			 50.9Tailing
			 
			 64.05
			 
			 3.61
			 
			 1.08
			 
			 0.75
			 
			 11.2
			 
			 3.2
			 
			 13.6Head
			 
			 100.00
			 
			 20.70
			 
			 21.97
			 
			 3.53
			 
			 100.0
			 
			 100.0
			 
			 100.0Hypothetical example of a
				  combined Pb/Zn bulk concentrate35.95
			 
			 51.13
			 
			 59.16
			 
			 8.48
			 
			 88.8
			 
			 96.8
			 
			 86.4






33



---



The 80% passing size to flotation for this
		test was not accurately known, but it is believed to be somewhere between 100
		microns and 125 microns. Kennecotts concerns about liberation and the
		occurrence of middlings are apparent in the high recovery of zinc into the lead
		concentrate and the assays of silver and lead in the zinc concentrate, though
		it must be noted that the ratio of lead to zinc in the feed to this test was
		high at 6.2 : 1. This ratio, as in the processing of all polymetallic ores,
		governs the success or failure of maximizing recovery of zinc into the zinc
		concentrate at an acceptable concentrate grade, i.e., 49% Zn or higher. In
		other words, if the ratio was lower, 1 : 1, or the reverse, the probability of
		optimizing the economic production and the quantity of a saleable zinc
		concentrate would be enhanced. A hypothetical example of this is shown below by
		increasing the zinc head grade to 6.90% and reducing the assay of silver and
		lead in the zinc concentrate, e.g.

SEPARATE Pb/Zn CONCENTRATES

(an example)

(Pb/Zn Ratio 3:1)

TABLE NO.6



Assay
			 
			 Recovery, %Product
			 Wt %
			 
			 Ag oz/ton
			 
			 Pb %
			 
			 Zn %
			 
			 Ag
			 
			 Pb
			 
			 ZnPb Concentrate
			 28.93
			 
			 61.03
			 
			 65.10
			 
			 8.47
			 
			 85.3
			 
			 90.1
			 
			 35.5Zn Concentrate
			 7.02
			 
			 10.32
			 
			 19.95
			 
			 50.00
			 
			 3.5
			 
			 6.7
			 
			 50.9Tailing
			 64.05
			 
			 3.62
			 
			 1.04
			 
			 1.46
			 
			 11.2
			 
			 3.2
			 
			 13.6Head
			 100.00
			 
			 20.70
			 
			 20.90
			 
			 6.90
			 
			 100.0
			 
			 100.0
			 
			 100. 0Hypothetical example of a
				  combined Pb/Zn bulk concentrate35.45
			 
			 51.85
			 
			 60.00
			 
			 16.82
			 
			 88.8
			 
			 96.8
			 
			 86.4




In both cases, the effect of combining the
		lead and zinc concentrates into a bulk concentrate for sale, results in high
		assays of payable silver and lead (and gold) at higher recoveries compared to
		differential concentrates, but at the sacrifice of non-payable zinc.

Previous estimates of ore reserves which
		have been commissioned by Chief have been based on the production of separate
		concentrates for lead and zinc by differential flotation. In this Report, it is
		considered more practical and economical to produce a bulk concentrate as
		described above from ore zones which contain high ratios of lead to
		zinc.

A review of drill core assays supplied by
		Pincock, Allen, and Holt shows that, in addition to the majority of intervals
		which report high ratios of lead to zinc, there is probably a zone in the
		North-east part of the Burgin Mine in which the ratio of lead to zinc is most
		probably 1 : 1 to 1 : 3 in which case, production of a distinct zinc
		concentrate would be considered feasible but subject to definition of the
		desirable mineable tonnage in this zone.(See Figure No.5)





34



---



Accordingly, and for the Scoping Study
		stage, a program of metallurgical testwork should be designed based on access
		to new drill core which would intersect the Burgin extension zones of
		mineralization down dip, across strike, and below the water table from the
		cessation of mining by Kennecott. A preliminary review of possible drilling
		sites has been undertaken with Mr. David Jenkins, based upon existing
		drill stations and presumed access to them. The following duplicate holes are
		recommended:(See also Table
		No.4)

RECOMMENDED DRILL PROGRAM

TABLE NO.7



Drill Hole Twin
			 
			 
			 
			 Mineral ZoneSB -14
			 
			 Vertical
			 
			 High PbB-122
			 
			 Vertical
			 
			 High ZnB-68
			 
			 Vertical
			 
			 High PbCB-9
			 
			 Vertical
			 
			 High ZnCB-30
			 
			 Inclined
			 
			 High PbTo SETwo
			 
			 Vertical
			 
			 PotholeTwo
			 
			 Vertical
			 
			 from 1450 level


On the basis of selecting a minimum of three
		composites from each hole which will reflect a range of lead: zinc ratios and
		which could be combined across holes for similarity in ratio based upon
		mineralogical examination, the following preliminary scope of metallurgical
		testing is proposed:


			 Conduct mineralogical examination to
				  define mineral liberation characteristics;



			 Conduct Bond Work Indices for
				  determination of specific power requirements, as well as Bond Abrasion Indices,
				  for single-stage ball milling;



			 Investigate optimum grind to
				  flotation for a bulk concentrate and extend to differential flotation as
				  considered appropriate;



			 Conduct rougher flotation tests to
				  determine optimum flotation times and reagent conditions;



			 Investigate regrind tests and
				  preliminary cleaner tests, supplemented by mineralogical examination of
				  products;



			 Conduct final batch-scale rougher,
				  regrind, and cleaner tests with differential flotation as appropriate;



			 Conduct thickening tests on bulk and
				  differential concentrates as appropriate;






35



---




			 Conduct filtering tests on bulk and
				  differential concentrates as appropriate;



			 Concentrate quality analysis.


Following an enquiry to an internationally
		recognized metallurgical testing laboratory, the approximate cost of such a
		program is estimated to be in the range ofUS$ 125,000toUS$ 150,000.





36



---



BURGIN
		WATER

The Extension ore-body must be adequately
		dewatered, before commencing mining operations. This requires the establishment
		of an effective mitigation development strategy and provision of the necessary
		equipment. The first stage in the activities must be a thorough technical
		review of the available data, hydrology testing from the proposed Burgin drill
		program and the definition of the work required to produce optimum results. The
		eventual solution can be addressed from two broad points of view.


			 The disposal of water to facilitate
				  mining



			 The use of the water to produce an
				  economic benefit


Prior to the results from the Burgin Mine
		work program, we are suggesting that it is only necessary to pursue the water
		Application and Sales segments of the desalination plan. The cost to establish
		hydrologic conditions in the mine are covered in the mine program. The budgeted
		amount to forward all other water activities in the time frame under
		consideration here (nine months), is $30,000.

HYDROLOGY

The hydro-geologic conditions, relative to
		the Burgin Mine, are described by Morris (1964), Lovering and Morris (1965) and
		Morris and Lovering (1974). Hydrogeologic conditions in the Goshen Valley
		immediately east of the mine have been described by Bissell (1963) Cordova
		(1970) and Dustin and Merrill (1980).

The Burgin Extension ore-body is highly
		fractured, with groundwater entering the mine workings via faults and
		fractures. Data by Lovering and Morris (1963) indicate that groundwater in the
		vicinity of the mine flows eastward to the Goshen Valley - through the Tertiary
		Packard Quartz Latite. Lithologic data collected by Sunshine Mining Company
		from monitoring wells located approximately 2,200 ft northeast of the old
		Kennecott disposal ponds indicate that, due to a flatter groundwater gradient
		than stratigraphic dip, groundwater flows through the Tertiary Pinyon Creek
		Conglomerate. This unit overlies the Packard Quartz Latite. The Packard Quartz
		Latite is predominantly a medium-grained quartz latite porphyry. It is
		generally divisible into an upper unit of vitrophyre and tuff with a thickness
		of up to 500 feet in the region, a middle unit of quartz latite porphyry that
		reaches a local thickness in excess of 2,700 feet, a lower unit of vitrophyre
		as much as 200 feet thick, and a basal unit of fine-grained tuff as much as 700
		feet thick (Morris and Lovering, 1979).

The Pinyon Creek Conglomerate is a poorly
		sorted conglomerate consisting of boulders and cobbles in a matrix of grit and
		sand. This formation reached a thickness of at least 1,000 feet in some
		locations of the region (Morris and Lovering, 1979).

A test hole (DP-1) was drilled by Sunshine
		immediately east of the disposal ponds in 1986 as part of a previous
		hydrogeologic investigation (Earthfax Engineering, 1986). This hole encountered
		coarse-grained alluvial sands and gravels in a silty matrix to a depth of 117
		feet, at which point the Pinyon Creek Conglomerate was encountered. The





37



---



hole was advanced to a depth of 190 feet
		within the conglomerate and abandoned due to drilling problems (inadequate rig
		size) prior to encountering groundwater. The relatively shallow depth at which
		bedrock was encountered beneath the ponds (compared with the thickness of the
		unconsolidated valley fill in Goshen Valley to the east) suggested that the
		primary Basin and Range fault that resulted in the formation of the east Tintic
		Mountains and the adjacent Goshen Valley exists east of the ponds and that the
		ponds overlie a pediment bedrock surface.

Two additional holes (DP-2 and DP-#) were
		drilled by Sunshine in 1986 and completed as monitoring wells approximately
		2200 feet northeast of DP-1. These holes encountered groundwater in the Pinyon
		Creek Conglomerate at a depth of approximately 340 feet. Static water levels in
		the wells rose to about 270 feet below ground surface to an elevation of
		approximately 4,536 feet above mean sea level (amsl). With a water level in the
		old Burgin Mine of approximately 4,550 amsl (see Lovering and Morris), 1965)
		the hydraulic gradient between the mine and the two existing monitoring wells
		east of the disposal ponds is about 3.4 feet per mile.

Hydrogeologic conditions beneath the stream
		channel leading to the disposal ponds are unknown. However, based on a review
		of the published geologic reports and surface exposures, it is expected that
		the alluvial deposits and the Pinyon Creek Conglomerate thin toward the west.
		The uppermost saturated zone occurs in the Packard Quartz Latite in much of the
		area west of the alluvial fan on which the disposal ponds are located.

Cordova (1970) indicates that groundwater
		occurs in Goshen Valley in four aquifers. In descending order, these aquifers
		are a water table aquifer, a shallow Pleistocene aquifer (existing under both
		confined and unconfined conditions), a deep Pleistocene aquifer (confined) and
		a Tertiary aquifer (confined). In the western portion of Goshen Valley (i.e.
		that portion of the valley closest to the disposal ponds), Cordova (1970)
		indicates that the water table aquifer and the shallow Pleistocene aquifer are
		interconnected. The direction of groundwater flow beneath western Goshen Valley
		is generally northeastward toward Utah lake.

Water that was encountered in the old Burgin
		Mine was in excess of 140° F and moderately saline (with total dissolved
		solids concentrations varying from about 6,000 to 8,000 milligrams per litre).
		Klauk and Davis (1984) and Cordova (1970) similarly identified groundwater of
		moderate salinity in Goshen Valley. The chemical characteristics, of the
		southern Goshen Valley groundwater, are different to groundwater in southern
		Utah Valley. Goshen Valley ground-waters have elevated temperatures as well as
		elevated concentrations of sodium and chloride. Many of the Goshen Valley
		ground-waters are chemically similar to water encountered in the old Burgin
		Mine as well as thermal groundwater discharging at Lincoln Point in Utah Lake
		(Klauk and Davis, 1984). Previous groundwater sampling in Goshen Valley has
		indicated that this similarity is naturally occurring, rather than being
		influenced by mine-water discharges (Hood, 1975; Dames & Moore, 1985).
		Nevertheless, this factor must be clearly understood and acceptable to all
		involved in the Burgin Water Application.





38



---



Davis and Cook (1983) found that Goshen
		Valley is a graben comprised of several structural blocks which are filled with
		sediment. The many faults associated with these blocks probably form conduits
		for upward migration of naturally-occurring thermal, saline groundwater, which
		would mix with good-quality groundwater at shallower depth. The occurrence of
		saline, thermal discharges at Lincoln Point suggests some upward migration of
		saline groundwater along faults.

MINE
		DEWATERING

During the production operation by Bear
		Creek at the Burgin Mine, the workings extended into the water table. Some of
		the events at the time, and the pumping problems encountered, have been
		described earlier in this report.

A pre-requisite to the compilation of an
		effective water removal plan is to define the quantities of water that need to
		be removed as mining extends further into the standing water table.

The ore-body Extension extends from the 1300
		level in the Burgin No. 2 shaft (4,395 el.) down to 3,939ft el., i.e. a
		vertical distance of some 455 ft. The current elevation of the water, i.e. the
		water table, is 4,563 ft el., hence it will be necessary to dewater at least
		624 ft of vertical distance.

As an aid to future technical investigations
		and relative to the quantity of water to be removed, the following historic
		facts are available. The mine is subjected to extensive flows of warm corrosive
		saline water containing elevated levels of arsenic, lead and total dissolved
		solids (TDS). The TDS on average will be 7,000 mg/litre.


			 The provision of adequate wells and
				  other dewatering components must be an absolute priority. In the past this was
				  neglected and the workings were not generally water free.



			 The eastern end of the ore-body
				  produces considerably less water than the west.



			 By the end of 1967 with the bulk of
				  production above the water table, the mine was pumping 4,400 U.S. gpm via two
				  wells, one from surface and a 17.5 Ø well from the 1200 ft
				  level.



			 By 1969, several churn wells were in
				  operation from the mine workings and pumping was at 5,200 U.S. gpm by year end.
				  Mine workings were still above the 1200 level and concentrated on 1050
				  level.



			 During 1970 the 1300 level was
				  opened and drifting commenced. Production was above the 1200 level. During the
				  year the drainage peaked at 10,000 U.S. gpm of 150°F water but by year end,
				  had reduced to 9,000 U.S. gpm. It stayed at this level until closure in 1978.
				  Mine production increasingly was






39



---



concentrated off the 1300 level and many of
		the stopes were still wet creating very poor ground conditions with resultant
		mining inefficiencies.

In a study completed by Sunshine Mining
		Company in the 1980s a relatively arbitrary figure of 12,400 U.S. gpm was
		utilized for the quantity to effectively dewater the extension ore-body. This
		estimate was based upon the expected open surface area of the proposed mining
		operation. This figure is still the one quoted in the more recent data, and has
		been used in projected development plans.

Dames and Moore, carried out an office study
		in 1985. They were not so certain 12,000 U.S. gpm had validity and opined that
		flows could theoretically reach 40,000 U.S. gpm. Chester Engineers, a
		subsidiary of U.S. Filter Operating Services, did a limited hydrology review
		from geological interpretations and estimated between 17,000  22,500 U.S.
		gpm may be more appropriate.

Over the years, there have been several
		investigations and studies into methods of resolving the water issue. In 1995
		ESA Consultants of Ft. Collins, Co., utilizing all the previous work, carried
		out what they termed a Screening Study on behalf of Chief. A
		summary of the available solutions is summarized below:

POSSIBLE WATER SOLUTION

TABLE NO.8



METHOD
			 
			 CAPITAL COST
			 
			 OPERATING COST/YEAR1. On site water treatment to public
				  quality standards. Brine evaporated in surface ponds.
			 
			 
			 $
			 42.5MM
			 
			 
			 
			 $
			 8.0MM2. On site water treatment to public
				  quality standards. Brine disposal underground.
			 
			 
			 $
			 36.2MM
			 
			 
			 
			 $
			 8.5MM3. Solar evaporation in the lower
				  Sevier River south of Deseret.
			 
			 
			 $
			 44MM
			 
			 
			 
			 $
			 4.5MM4. Solar evaporation in the Crater
				  Springs Area north of Delta.
			 
			 
			 $
			 46.6MM
			 
			 
			 
			 $
			 4.4MM5. Solar evaporation north of Cedar
				  Valley.
			 
			 
			 $
			 24.2MM
			 
			 
			 
			 $
			 2.2MM6. Subsurface disposal into the
				  Chief Mine.
			 
			 
			 $
			 13.5MM
			 
			 
			 
			 $
			 3.7MM7. Subsurface disposal into the
				  Tintic Standard Mine.
			 
			 
			 $
			 7.2MM
			 
			 
			 
			 $
			 1.1MM


Notes:

1)
			 Above estimates are in 1995 US
				  dollars.


2)
			 Study assumes treating 12,000
				  gallons per minute.






40



---



Comments

At this stage of the Burgin development, the
		economics would lead to the conclusion that solutions (1) (2) or (7) are
		preferred. (1) or (2) are predicated upon economics. If the salt can be sold
		then (1) is the best option if not (2) would be preferable. Therefore, we
		conclude that the salt sales potential needs a speedy resolution and subject to
		the results, the water only treatment solution along with brine disposal into
		an adjacent mine, be the only desalination concept developed further. If the
		desalination concept is dropped, then efforts for disposal should concentrate
		on an environmentally acceptable old mine disposal system. Details of the
		desalination possibilities are described in more detail later on in this
		section of the report.

There has been no detailed hydrological
		evaluation to accurately determine the geothermal water conditions that will
		prevail during the mining of the Burgin extension ore-body. Sunshine did
		propose an extensive drawdown test which was never implemented. This test
		consisted of two large surface production wells, together with four smaller
		monitoring test wells. The main wells, which would subsequently be utilized for
		mine drainage, would extend 300 ft below the ore-body. The smaller monitoring
		wells would be drilled to 1,800 ft from surface. The testing program was to
		continue for a 12  18 month period.

There were two pumping set-ups under
		consideration.

(a)
			 Heavy duty, deep well pumps capable
				  of handling hot brine solution. Their use would depend upon the results from
				  the two wells used for the hydrology tests. Obviously more wells would be
				  needed to cover the perimeter of the ore-body. Numbers between 5  8 wells
				  have been suggested, the final number will be partly defined by the test
				  program and fully determined by actual application.


(b)
			 The alternate to the surface wells
				  would be a series of lighter duty wells installed around the ore-body from the
				  Burgin 1050 level. These pumps would discharge to a main pumping level. This
				  level would be equipped with sumps and bulkheads along with high-head pumps
				  boosting the water to surface.


WATER
		APPROPRIATION

The water basin adjacent to the Burgin Mine
		is the Utah Lake Water District. This district is fed primarily by Utah Lake,
		which was closed from new applications on November 15th, 1995. Utah
		Valley extends from Point of the Mountain in the north to Santaquin Ridge to
		the south. The two locations are approximately 32 miles apart. The Valley is
		bordered to the east by the Wasatch Mountains and a series of low lying hills
		on the west. Utah lake lies in the middle of the District and is some 26 miles
		long (N-S) and 12 miles wide (E-W). The lake is fed by a series of streams
		flowing from the Wasatch Mountains. The northernmost is Dry Creek which flows
		through Alpine Canyon.





41



---



Proceeding southerly the other
		streams/rivers are American Fork, Provo River, Hobble Creek, Spanish Fork
		River, Peternect Creek and Summit Creek which flows out of Santaquin Canyon.
		Water also enters the lake via Current Creek which flows from the Mora
		Reservoir near Nephi. These waterways are the only significant source of water
		in Utah Valley they recharge the lake and aquifers under the Utah
		Valley.

As part of the past desalination efforts, on
		September 4th, 1998, Chief, on behalf of its wholly owned subsidiary
		Tintic Utah Metals LLC, filed an Application with the Office of the State
		Engineer (Utah) to appropriate 29,000 acre  ft of water per year. This
		water is located below the water table (4,563 ft el.) in the old Burgin Mine.
		On March 30th, 1999 the State Engineer held a public hearing on the
		Application to Appropriate (A71750). At the hearing Chief presented
		expert and written testimony, which remains of record. Various
		Protestants to the Application, particularly the Central Utah Water Conservancy
		District (The District) requested time to review the materials submitted by
		Chief and to submit a response to those materials. The Engineer agreed and gave
		the Protestants 30 days to respond and Chief a consecutive 30 days to reply.
		All the Protestants had interest in and concerns about the Applications
		impact on the local Utah Lake/Goshen valley water balance.

Eventually, on May 27th 1999
		after a series of granted delays, both parties had completed independent
		studies. Of the nine initial Protestants, only The District had prepared
		comments to the Chief filings. Responses are summarized as follows:


			 The District agreed that the Burgin
				  geothermal system was very deep and large, four to seven miles below the Goshen
				  Valley; it contained as much as 118 million acre ft of water and was not
				  associated with local precipitations or run-off.



			 The Protestant was concerned that
				  this deep seated system communicated with and was tributary to the deep basin
				  fill aquifers in the Utah and Goshen valleys. The Protestants however did not
				  have data to validate such a hypothesis. Chief presented argument from
				  acknowledged Independent Expert Consultants that there was no
				  connection. One of the reasons given was the fact that during 11 years of
				  pumping the water from the mine, (1967  1978) there were no changes at
				  all to the Goshen Valley reservoirs and aquifers. Another of the given reasons
				  was that the wells in the Goshen Valley all contain better quality water with
				  depth. Utah Lake has been in existence approximately 10,000 years and
				  associated Goshen Valley aquifers are some 2,000 ft thick. If, over this time,
				  a connection was present then, it was opined, the deeper one goes into the
				  aquifer the worse the water quality should be.



			 The Protestant argued correctly that
				  there was no un-appropriated water in the Utah Lake/Goshen Valley area and
				  assumed that the Burgin Application was part of the area. Chief responded with
				  sound technical, hydro-geological and practical counter arguments.






42



---








			 The point was made, by the
				  Protestant, that the plan presented for the utilization of the water by Chief,
				  did not unequivocally demonstrate physical or economical feasibility. Chief
				  responded that these aspects were not within the jurisdiction of the State.
				  Chief, its shareholders and bankers would decide as to the viability of the
				  project once a final mine feasibility study was completed. This could not
				  happen until the resolution of the water problem was clear and economically
				  defined.


In August 1999, the State Engineer denied
		the Application but he left the door open for a new Application.
		Thus, on February 15th, 2000 Chief submitted a new Application. This
		Application modified the original, such that Chief was to install and monitor
		two wells at Lincoln Point and Goshen Hot Springs, and was to come to an
		Understanding with the Protestants.

On September 14th 2000 a further
		hearing was held before the office of the Engineer. At this hearing each of the
		Protestants and Chief presented their respective arguments, facts and positions
		in favor of or against the granting of the Application. Following this meeting,
		as a result of negotiations, the Engineer was requested to delay a decision
		until Chief and the Protestants could agree a formula by which they could all
		approve the granting of the Application. The Engineer agreed to the
		delay.

The situation today (October 2005) is that
		there have been several meetings between Chief and the Protestant and a draft
		Settlement Agreement tabled on October 18th 2001. Finalization of
		this Agreement is still outstanding.

DESALINATION

Summary

Under the conditions and assumptions used,
		published data support costs for 75% volume yield desalination of Burgin
		mine-water:


			 Capital $24 million



			 Operations and Management
				  $0.29/m3 ($348/ac-ft)


Critical conditions and assumptions
		include:


			 Exclusion of all costs relating to
				  raw water supply to the plant, retentate (brine) disposal and distribution of
				  desalinated water past the plant battery limit.



			 No mixing of brine (disposed by
				  injection to abandoned mine workings) with influent to the Burgin Mine
				  dewatering system, i.e. no change in Minewater salinity over the project
				  life.



			 Local power availability at
				  $0.05/kWh on a non-interruptible basis



			 No allowances for permitting,
				  negotiation of water rights and/or any fees associated with the water supply
				  and brine disposal






43



---



The indicated costs allow a margin for
		profitable operation if product water, nominally 400ppm TDS, can be sold for
		$600-700/ac-ft.

Note that no project specific contingency
		allowances have been applied to the costs presented above.

If other project factors are favourable, the
		next stage of evaluation should include:


			 Hydrogeological review of
				  underground injection brine disposal to identify and quantify any effects on
				  salinity of Burgin Mine dewatering flows



			 Confirmation of power
				  availability/cost



			 Development of preliminary site
				  layout and general design criteria


Satisfactory results from these studies will
		provide a basis for approaching Reverse Osmosis system vendors to obtain
		site-specific budget capital and operating cost estimates for
		desalination.

Background

Reopening of the Burgin Mine to exploit
		deeper resources will require a significant dewatering program and a practical,
		cost-effective method of disposal of the product water.

Estimates of the volume flow required for
		mine dewatering range upward from 12,000USgpm. This value is used as the base
		case for the current study, corresponding to a flow of 45.4 cubic meters/minute
		(65,376m3/d, 23.9 million m3/y or roughly 20,000 acre
		feet (ac-ft) per year, of warm (140F) brackish (7000mg/L total dissolved
		solids, TDS) water from below the planned development.

The base case for Mine-water disposal is
		currently anticipated to be discharge into the (abandoned) workings of one or
		more adjacent mines, with the expectation that the water will permeate back to
		the deep geothermal source aquifer without significant short-circuiting into
		the Burgin Mine dewatering zone.

This assumption is also basic to concepts
		assessed in the current study. It must be noted that independent assessment(s)
		will be required to confirm both the hydro-geological feasibility and permit
		acceptability of the concept.

A previous study, (U.S. Filter Study) has
		been reviewed. Options from this study included:


			 Evaporative pond disposal. This was
				  summarily rejected on the basis of large pond area and cost.



			 Forced evaporation pond(s). This was
				  also summarily rejected on energy cost as well as pond area/cost.






44



---




			 High yield desalination by a
				  sequence of reverse osmosis (RO) and evaporative (mechanical vapour
				  compression, MVC) technologies followed by spray evaporation.
				  Ultimate salt disposal was as bulk crystal/saturated brine in a lined surface
				  impoundment.


The latter option was examined in some
		detail, resulting in an order of magnitude treatment capital cost estimate of
		$150 million or $2550/m3/d. This is above the high end of the range
		of specific capital costs attributed to the US Army Corps of Engineers for RO
		seawater desalination. Such a high capital cost is plausible based on the
		thermal (evaporative) component and high brine disposal costs. However, it
		implies a requirement for high revenue from purified water to achieve
		break-even or profitable operation.

The U.S. Filters report cites regional bulk
		water pricing in the range of $1,500-$1,900/ac-ft. More recent investigation
		suggests that the current local value of bulk desalinated water is likely to be
		in the $600-700/ac-ft range.

The objective of this study is to review
		alternative desalination strategies which might result in cost-effective
		processing of Burgin Mine water in the context (see above) of practical
		disposal of concentrated brine by injection to local abandoned mine workings
		without significant effect on salinity of the Burgin Mine dewatering
		flow.

Scope and Battery
		Limits

The scope includes a raw water
		cooling/storage/aeration pond and reverse osmosis (RO) desalination
		facilities.

Battery limits are from discharge of the
		mine water to the process feed pond to the intake of the potable water
		discharge from storage.

Exclusions

Mine dewatering to the point of discharge to
		the raw water pond and potable water discharge (to pipeline or aquifer) from
		the potable water storage are specifically excluded.

Overview and
		Technology Selection

The 7000 ppm salt content of Burgin
		mine-water is significantly above the level at which ion exchange and
		electrodialysis are practical. This limits the potentially applicable
		technologies to reverse osmosis and evaporative technologies.

For relatively low raw water salinity (i.e.
		brackish water), RO costs are typically significantly lower than evaporative
		methods. Therefore, for the purpose of the immediate future work, relative to
		the water issue, Reverse Osmosis is the selected desalination route. Since we
		are suggesting that the program during the next nine months be restricted to
		the Appropriation and possible water sales, any further detail into
		desalination improvements can be safely delayed.





45



---



For example, the possibility of combining a
		low yield MVC process with hot softening and cooling of raw water to the
		(about) 100F target for RO feed might be an appropriate trade-off study at the
		feasibility level.

Overall Treatment
		Concept

The conceptual treatment scheme for
		assessment is:


			 Raw water cooling with chemical
				  pretreatment. Precipitates and suspended solids are disposed to a surface
				  impoundment.



			 Single stage RO to about 75%(v/v)
				  recovery of desalinated water (permeate) at about 400ppm TDS.



			 Retentate (brine, 27-28,000ppm TDS)
				  to underground disposal in non-connective local abandoned mine workings.


Pretreatment of the entire volume of raw
		water is conceptually preferred so that solids and scale will not block
		percolation of brine into the re-injection formations.

Reverse Osmosis 
		Process Considerations

For the Burgin Mine, raw water processing
		requires pre-treatment including;


			 Cooling from 140F to <100F



			 Oxidation and hydrolysis of the
				  (minor) ferrous iron and manganese (II) components



			 Separation of suspended
				  solids



			 Rejection of divalent ions (calcium,
				  magnesium, sulphate) and silica



			 Polishing filtration



			 pH adjustment



			 Anti-scalant addition


Treatment chemicals represent a significant
		process cost and the solid products (sludge) require surface impoundment. As
		noted in the previous section, raw water suspended solids and precipitates must
		be disposed of on surface to optimize infiltration of waste brine into the
		re-injection formations. Surface disposal will require a lined impoundment due
		to the salt content of the aqueous phase. The slurry for disposal will be
		alkaline and non-hazardous.

Specification of the actual suite of
		chemicals for water pre-treatment will require testwork optimization for both
		performance and cost-efficiency.

For RO, a single stage system with phased
		membrane replacement to maintain average permeate salinity at or below the
		nominal 400 ppm TDS target and retentate (reject) salinity in the 28,000 range
		will result in a volumetric desalinated water yield in the 75% range, i.e.
		9,000 U.S. gpm or 14,900 ac-ft/year.





46



---



If future market studies indicate a premium
		(in value or marketability) for lower salinity product, the extra capital cost
		of a two stage RO system could be justified due to better operability as well
		as product quality.

Treatment
		Costs

For pre- or final feasibility assessments,
		the usual and appropriate approach of costing RO water treatment is to develop
		design criteria and/or a performance specification as a basis for obtaining
		vendors budget quotations. Insufficient data are available for this
		approach at present.

A computer search has been conducted to
		identify and collate published data and estimates for a range of analogous
		water treatment projects. These data, with allowance for site specific costs
		have been used to identify a probable range of total (capital + operating) unit
		costs as well as capital and operating cost estimates for desalination of
		Burgin mine-water.

Cost
		Factors

The principal reference(Ettourney, H.M., El-Dessouky, H.T., Faibish, R.T., et
		al Evaluating the Economics of Desalination, Chemical Engineering
		Progress, pp 32  39, December 2002)for this section presents an overview of process cost
		impacts and a simplified overall cost estimation model. Identified major cost
		factors for RO processing include:


			 Raw water salinity. Lower salinity
				  allows higher volumetric recovery with lower specific (i.e. $/unit) power and
				  chemical consumption as well as less scale related downtime.



			 Plant capacity. Larger plants
				  provide economy of scale.



			 Qualified labour. This has a major
				  effect on availability and capacity.



			 Energy cost



			 Plant life and amortization
				  arrangements


The Burgin site appears favourable (i.e.
		indicating low range costs) in terms of raw water salinity (low relative to
		seawater), plant capacity (high enough to be insensitive to size) and location
		with respect to availability of qualified labour.

In the case of the Burgin Mine, desalination
		plant life would have to be significantly longer than the mine life indicated
		by presently identified resources; normal plant lifetimes are in the 20-30 year
		range. Negotiation of suitably long water tenure would be an essential
		pre-requisite to establishing such a facility.





47



---



Direct Capital
		Costs

Direct capital cost components of on RO
		plant normally include:


			 Land



			 Well or intake construction



			 Process equipment and membrane
				  modules



			 Auxiliary equipment



			 Building(s)


For the purpose of comparison with other
		projects:


			 Land for the RO facility is assumed
				  to be available as part of the mine development package



			 Capital cost of the wells (mine
				  dewatering) are  for the purpose of this study  attributed to the
				  mine, although this could be split depending on results of overall project
				  financial analysis.



			 Auxiliary equipment can only be
				  specified as part of a development plan and identified market for product
				  water. Previous study data suggest connection costs (20,000ft of buried
				  pipeline plus a groundwater injection system) are likely to be above
				  normal.



			 Buildings suitable for high
				  elevation winter conditions may be higher cost than for lower elevation coastal
				  conditions.



			 Membrane costs in the range of
				  $500-$1,000 per 50m3/d module.


Indirect Capital
		Costs

Capital indirects, as for other large
		projects, include freight and insurance, construction overheads and owners
		costs.

For the Burgin Mine, the mid-continent
		location with existing road access, close to major transport arteries is
		favourable.

Contingency

Contingency allowances have not been added
		to the estimates presented in this study. Although it is likely (but cannot be
		confirmed) that some of the data used already contain contingency allowances,
		the author considers that it would be prudent to add a significant contingency
		allowance for site specific factors in any financial analysis using these
		data.

Operating
		Costs

Principal elements of RO plant operating
		costs are:


			 Electric power



			 Labour



			 Membrane replacement






48



---




			 Maintenance and spares



			 Chemicals



			 Insurance


Specific power cost and availability for the
		Burgin site have not been provided, so $0.05/kWh is assumed. Continental USA
		power costs are at the low end of the reported $0.04-0.09/kWh spectrum of
		worldwide RO plant power costs.

Labour costs are expected to be mid-range
		for USA locations.

Membrane replacement is highly variable
		(5-20%/y of initial membrane cost) depending on raw water and operator
		efficiency. For Burgin, brackish water and (assumed) high quality operating
		labour suggest a low range, 7.5%/y replacement allowance.

Chemical costs are typically in the
		$0.03-0.05/m3 range. The raw water is low in most problem species,
		so a low end $0.03 allowance is considered appropriate.

Project Rationale

Under the assumption that the Burgin
		Mine-water could be pumped untreated to abandoned mine workings for disposal,
		possible justifications for desalination are:


			 Reduced cost of water disposal due
				  to reduction (75%) in volume pumped to final disposal sites and/or better or
				  more assured percolation of the lower reject volume.



			 Stand-alone project economics of
				  desalination


Potential benefits of a reduced volume for
		disposal cannot be quantified at present, so the only strong justification for
		water desalination would be acceptable stand-alone project economics.

In the context of currently anticipated
		revenue ($600-700/ac-ft), all-in project costs including capital payback and
		operations should not exceed $0.50/m3 of product water
		($600/ac-ft).

For reference purposes, there are frequent
		literature references to $0.50/m3 as a near term achievable target
		for seawater desalination. Since coastal seawater is much higher (30-35,000ppm)
		salinity than Burgin raw water and even the salinity of reject from the
		conceptual RO (28,000ppm) is below that of seawater, it is not unreasonable to
		expect that this target may be met.

All-in Cost
		Comparators

Data on estimated and (reported) actual
		all-in costs for a range of RO desalination projects are summarized
		below:





49



---



COMBINED CAPITAL + OPERATING
		COSTS

TABLE NO.9



Capacity, m3/d
				  Product
			 
			 Cost,$/m3
			 
			 Reference and Notes20,000
			 
			 0.64
			 
			 Single stage seawater, 1987
				  data20,000
			 
			 0.76
			 
			 Two stage, as above94,625
			 
			 0.55
			 
			 Florida seawater, 1999Unspecified
			 
			 0.37
			 
			 California, brackish estimateUnspecified
			 
			 0.28-0.41
			 
			 Florida, brackish estimateUnspecified
			 
			 0.40
			 
			 Texas, brackish, estimateUnspecified
			 
			 <0.50
			 
			 FAO, general estimate


The 1987 values are old, reflecting out of
		date membrane costs, and the low end Florida brackish estimate is believed to
		relate to very low salinity feed. This leaves a range of all-in costs
		$0.37-0.55/m3. If the high value in this set is rejected on the
		basis that it is for seawater, the range is $0.37-0.50 ($444-600/ac-ft), which
		is within the target range for achieving some net project revenue from water
		treatment with a minimum water revenue of $600/ac-ft.

Capital
		Cost

Only one reference was found giving a range
		of capital costs ($380-$560/m3/day) attributed to the US army Corps of
		Engineers for Florida locations(Desalination by reverse osmosis
		http://www.ors.org/osde/publications/unit/oea59e/ch20.htm). A mid range estimate of $460 indicates an order of
		magnitude capital cost of $23.9 million.

Operating
		Cost

Direct plant operating costs are estimated
		below based onEttourney et
		alexcept for membrane costs which are
		reported, recent, (2003) actual costs for a seawater installation in Israel(Glueckgtern, Pinhas Desalination;
		Current Situation and Future Prospectus; The Begin-Sadat (BESA) Center
		for Strategic Studies, Bar Ilam University, Israel, Water Article
		No.1). This cost is at the low end of
		literature reported values, but does reflect the widely reported significant
		improvements in membrane performance and reductions in cost over the past two
		decades.





50



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DESALINATION OPERATING COSTS

TABLE NO.10



O&M Item
			 
			 $/m3Membrane replacement
			 
			 0.035Chemicals
			 
			 0.03Labour
			 
			 0.05Maintenance materials
			 
			 0.02Insurance
			 
			 0.005Subtotal
			 
			 0.14Power
			 
			 0.15
			 (3kWh/m3 @
				  $0.05/kWh)Total
			 
			 0.29


It should be noted that the operating cost
		is very sensitive to electric power cost and consumption. The 3kWh/m3
		consumption is based on low salinity and is in the bottom quartile of power
		requirements for RO facilities. Also, the total operating cost is at the low
		end of the US Army Corps of Engineers estimate ($0.28-0.41/m3) for
		brackish water desalination at Florida sites.

As a check, total operating costs typically
		account for about 60% of all-in costs, which is at least plausible for a
		capital cost in the $24 million range.

Conclusions and
		Recommendations

Under the conditions and assumptions used,
		published data support costs for 75% volume yield desalination of Burgin
		mine-water:


			 Capital $24 million



			 Operations and Management $0.29/m3
				  ($348/ac-ft)






51



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Critical conditions and assumptions
		include:


			 Exclusion of all costs relating to
				  raw water supply to the plant, retentate (brine) disposal and distribution of
				  desalinated water past the plant battery limit.



			 No mixing of brine (disposed by
				  injection to abandoned mine workings) with influent to the Burgin Mine
				  dewatering system, i.e. no change in Minewater salinity over the project
				  life.



			 Local power availability at
				  $0.05/kWh on a non-interruptible basis



			 No allowances for permitting,
				  negotiation of water rights and/or any fees associated with the water supply
				  and brine disposal


The indicated costs allow a margin for
		profitable operation if product water, nominally 400ppm TDS, can be sold for
		$600-700/ac-ft.

Note that no project specific contingency
		allowances have been applied to the costs presented above.

If other project factors are favourable, the
		next stage of evaluation should include:


			 Hydrogeological review of
				  underground injection brine disposal to identify and quantify any effects on
				  salinity of Burgin Mine dewatering flows



			 Confirmation of power
				  availability/cost



			 Development of preliminary site
				  layout and general design criteria


Satisfactory results from these studies will
		provide a basis for approaching RO system vendors to obtain site-specific
		budget capital and operating cost estimates for desalination.

POTENTIAL WATER
		SALES

Based on current information, it is believed
		that the hydrology of the Wasatch Front is well understood. The annual total
		flow to Utah Lake and the recharge of the valley aquifers is believed to be a
		known finite quantity. All of this water is believed to have been appropriated
		and put to beneficial use. In short, under the present population and use
		conditions, there is no more water available. A recent study projects that the
		population on the Wasatch Front will rise from 1,200,000 to 5,000,000 in the
		next 50 years. The conclusion is obvious. There is a need for additional water
		resources and the price of water will rise. It should be noted that the State
		of Utah consumes 81% of its water in the Irrigation sector, 14% for culinary or
		domestic use, and the remaining 5% is used in all other areas i.e. power
		plants, mining, industrial and aquaculture.

A recent water price check was made in the
		cities adjacent to the proposed Burgin Mine. Based upon a single family of 4,
		using an average 30,000 gallons per month i.e.1 acre ft per year, the following
		table shows the current prices being paid.





52



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CURRENT WATER COSTS

TABLE NO.11



Location
			 
			 Price per acre ftSalt Lake City
			 
			 $424Murray City
			 
			 $489Lehi City
			 
			 $610Provo
			 
			 $554Sandy City
			 
			 $660Spanish Fork
			 
			 $698


An indication of the future requirements can
		be found in a recent report from the Utah Board of Water Resources. This report
		states that the Municipal and Industrial (M & I) consumption in the Jordan
		Lake  Utah Lake area is 134,000 acre ft per year. Projections for the
		year 2020 rise to 207,000 acre ft/year, and for 2050, 338,000 acre ft/year. The
		same report states that there is only 50,000 acre ft of developable water in
		the area.





53



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ZUMA CLAY
		PROJECT

GENERAL

The Zuma property, which is part of the
		Chief ground is located in the East Tintic district. It lies in the
		southwestern part about 11⁄2 miles southwest of Dividend. This property
		consists of all or part of eight patented mining claims of approximately 125
		acres in total. Zuma has seen development in the past. A shaft 1,310 ft deep
		was sunk in 1922 and limited Au, Ag, Pb, Zn production was taken from
		underground.

During construction for an access road
		through the property in 1956, a halloysite clay deposit was exposed. Clays are
		formed by the mechanical and chemical breakdown of rocks (weathering). Clay
		deposits are classified as residual or sedimentary. Residual clays of which the
		Zuma deposit is an example, are deposited in the same location as they were
		formed. Clay resources are generally sub-divided into three groups.

(1)
			 bentonite and Fullers earth


(2)
			 kaolin, ball clay, halloysite and
				  refractory clay


(3)
			 Miscellaneous clay and shale


The second category is the area of interest
		in the clays found on the Zuma property. The halloysite clay consists of
		predominately endellite Al2 Si2O5
		(0H)4 2H20, which dehydrates, losing part of its water,
		to become halloysite. Both endellite and halloysite have been identified on the
		Zuma claims. The purest endellite is massive and dazzling white with a smooth
		lustrous surface resembling porcelain. Halloysite is dull chalky white and
		pulverulent. The Zuma potential for an industrial halloysite venture is
		demonstrated by the historic exploitation of the clays as well as recent
		sampling. A grab sample, taken by Consolidated Clay Materials Company in 1998,
		gave a grade of +80% halloysite. A second sample by a third party was reported
		to be close to 100%. In comparison the primary producer for ceramic use is New
		Zealand Clays Ltd., a subsidiary of Ceramco Ltd. Their in-place deposit grades
		about 80% halloysite.

RESOURCES

A lease contract was arranged between Chief
		and the U.S. Energy Corporation in 1972. In turn U.S. Energy negotiated a
		contract to sell halloysite clay to Filtrol Corporation. An estimated 50,000
		tons of halloysite was sold to Filtrol Corps plant near Salt Lake City
		where it was processed into a filter catalyst used in the cracking of crude
		oils into gasoline and other light petroleum derivatives. This production came
		from two adjacent open pits on the Zuma claims.

In 1990 Sunshine Mining Company drilled the
		clay deposits with a series of shallow holes. The Sunshine engineers
		subsequently estimated that there was a drill indicated resource of
		264,000 tons in the small area drilled off. Sunshine also sold 9,000 tons of
		clay material from the old dumps to Ash Grove Cement in Leamington, Utah for
		tests in their cement making process. The test was successful and all the low
		grade dump





54



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material was shipped during the period 1991
		 1993. Various articles and data viewed at the site indicate that the
		clay potential is equal to that of the adjacent Dragon Mine and upwards of
		1,000,000 tons are possible.

Our review of the historic data relative to
		a possible clay asset development, indicated that clays had been observed or
		may be geologically interpreted in the following areas covered by Chiefs
		claim block.


			 East Crown Point Consolidated
				  Propertywhich consists of parts of
				  eight patented mining claims totaling 92.324 acres. This property is contiguous
				  to the Zuma property and a portion of the Zuma clay deposit is located on the
				  Crown Point Extension Claim No.4.



			 The Iran King Property


Replacement clays have been observed.


			 The North Lily
				  Property


Halloysite clay has been observed.


			 Larsen clay pits.


HALLOYSITE

The mineral is an alumino-silicate clay that
		has special properties by virtue of having a common crystal form of a
		microtubule.





The composition of halloysite is A12Si205 (0H)4. It is relatively inert and
		has no reported toxicity or injurious characteristics. It occurs as an
		alteration product, hydrothermal or deuteric, of feldspars and is often
		associated as a mixture with kaolin, endellite, alunite,





55



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allophane and gibbsite. Silica in the form
		of quartz or adularia may be associated with these alteration products.

The crystal form appears to be a rolling of
		the typical plate like form of kaolin and some of its near relatives. It is
		close in composition to kaolin and included in the phyllosilicates
		(leaf or platey silicates). The principal distinction for halloysite is a
		higher water content and its crystal form as microtubules rather than laminae.
		It may be form by addition of water to the kaolin during alteration processes
		which addition would cause a distortion of the laminae to form the tubules as
		rolled laminae.

USES AND MARKET
		POTENTIAL

The high alumina content of halloysite makes
		it useful as an alumina source for cement production. As an example, there is
		an operation of a low quality halloysite in the Los Angeles Basin area, serving
		local cement plants. The pricing of such material is typically tied to freight
		rates for substituting alumina sources and a large part of any income is
		obtained from trucking as much as from the clay itself. The advantage of this
		market for the producer of high purity halloysite clay is that waste produced
		by selective mining for the high purity material can be sold and its production
		cost can be covered from such sales. Other lower quality/price markets include
		the brick and ceramics manufacturing industries, both of which exist in Utah,
		at no great distance from the Tintic Mining District.

A high purity halloysite product is used for
		fine porcelain, bone china and as a suspension agent in glazes. A high
		whiteness halloysite can find considerable value as a functional filler in a
		wider range of paints by virtue of the textural and strength giving properties
		of the tubule crystal structure. This acts as a mesh within the paint coat and
		can reduce the cracking induced by weathering, typical of most common paints.
		Industrial building paints (barn door and fence paints) do use some slightly
		coloured fillers as the pigment loading a so high as to cover up the colour
		derived from such lower grade filler.

The tubule shape of the crystals has also
		value as a large surface area for absorption and adsorption of liquids.
		Research by the U.S. Naval Research Laboratory has resulted in patents that
		allow the absorption of hydrophyllic and hydrophobic liquids (eg water
		solutions and oils respectively). This function is being used as a system
		making a slow release anti-fouling paint for ships and has potential in slow
		release applications for pesticides, fertilizers, fragrances and
		pharmaceuticals. A further filler use may include application to make air or
		inert gas voids in polymers for the production of strong, light weight
		materials.

This would appear to be a growth market that
		would be worth pursuing with a product of relatively high unit value that has
		unique physical properties due to the tubule shapes of the crystals.

The absorption ability of the microtubules
		also offers potential for its use as a relative low cost matrix for catalysts
		and as a specialty filter clay.





56



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OTHER SOURCES OF
		HALLOYSITE

Atlas Mining Company has recently purchased
		the Dragon Halloysite Mine in the Tintic Mining District (Juab County) Utah,
		and has announced an intention of production from a deposit of a reported
		8-900,000 tonnes of clay. It is not clear at this time how much of the deposit
		is a bright white clay as photographs on the Atlas web site
		(www.atlasmining.com/dragonmine.html) indicate that there is
		considerable contamination of iron oxides in the clay faces exposed for testing
		and future mining. Drill cores show a tan to mottled iron oxide colour in short
		lengths of boxed drill core. Some of the photos have been placed on the next
		pages that demonstrate the distribution of white clay within the mixed coloured
		minerals.





FIGURE NO.11

Atlas Mining Dragon Mine Operation





57



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The mixture of colours indicate that some
		processing will be required in order to obtain a high white high purity product
		to meet typical market requirement. The iron contamination is not a trivial
		problem for Atlas unless the photographs are misleading. The high quality
		kaolin deposits in Georgia and Cornwall are by contrast a blinding white
		colour, with very little contamination by other minerals.

Atlas claims that the only other production
		source at present is in New Zealand and it may be a source for the small high
		quality ceramic industry in that country. The supply from the North Star
		Minerals quarry in southern California that goes into cement production in the
		Los Angeles Basin is not officially qualified as halloysite and is actually a
		mixture of aluminous clays derived from the weathering and/or alteration of an
		anorthosite complex. It is unlikely that the operation would move into the
		higher priced markets but the operator would be a useful source of information
		for the low-end markets and may be a candidate as an operator or operations
		consultant for the subject clay deposit.

We have not looked into the trade and
		imports of halloysite other than the information available in Industrial
		Minerals Magazine and Mineral Facts and Problems, (1986) US Bureau of Mines
		publication. Unless the new markets have advanced more rapidly than expected
		and this remains unreported, the markets for halloysite will support a few
		producers provided the producers aggressively pursue those markets in the face
		of competing clays such as calcined kaolin.

ZUMA SHAFT AND LARSEN
		CLAY PITS

Other halloysite references for the Tintic
		Mining District show the Zuma shaft area and the Larsen clay pits in Juab
		County to include halloysite suitable for mineral sample material. A capsule
		description of this mineralization is located in www.mindat.org
		database. The minerals appear to be a mixture that would require typing in
		detail to determine the potential of the deposit. The size of the deposit has
		been reported to be similar to that of the Dragon Mine, just under 1 million
		tonnes. Atlas Mining indicates that they expect to sell their product at about
		US$400 per ton. High quality, high brightness, specialty kaolins sell at prices
		ranging from US$200 to US$350 per tonne and these are being used in some of the
		markets to which the halloysite can be applied. The special properties of the
		microtubules (nanostraws) of the halloysite may allow it to





58



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command a higher price in some markets. The
		further processing of the halloysite to add the active ingredients for slow
		release applications would allow capture of downstream value-adding and return
		a much larger net back to the operation.

Looking at the order of magnitude economics
		of an operation that would have a 20 year life, sufficient to encourage
		consumers that the product will be available for a reasonable long market life,
		we see the following scenario:



Deposit Size
			 
			 1,000,000 tonnes (tonnage assumed
				  to be developed)Life
			 
			 20 yearsMine type
			 
			 Open pit/quarryAnnual Mining Rate
			 
			 50,000 tonnesAnnual Gross Revenues*
			 
			 $
			 10,000,000Capital Cost
			 
			 $
			 2,500,000Operating Cost
				  (Annual)
			 
			 $
			 750,000Annual Net Revenues**
			 
			 $
			 8,000,000


|* | Using a mixed product price averaging
				$200 per tonne fob mine site. The freight profits for the 50,000 annual tonnes
				shipping are not considered here but this might offset much of the production
				cost.
|- | --------------------------------------------------------------------------------------------------------------------------------------------------------------------------------------------------------------------

|** | Assuming full sales of product.
|-- | -------------------------------

This range of costs and revenues should
		allow a payback of less than a year if full sales of production can be achieved
		from the start of operations. A typical industrial minerals operation has to
		grow its markets in the early years and smaller revenue streams are
		returned while stockpiling of product is required to support surges in sales
		that develop. For the 20 year life, we can expect a pre-tax net present value
		in the order of $100 to $150 million dollars.

The smaller sized operation has less
		environmental impact and is much easier to permit to production than a larger
		operation. Once permitted the expansion of operations is relatively simple.

RECOMMENDED WORK
		PROGRAM

The economic scenario outlined is a
		speculative assessment but indicates useful potential with a relatively low
		capital cost approach. A work program is required to confirm this potential.
		This work program would involve mapping and drill sampling of the deposit, with
		samples being sufficiently spread throughout the deposit to allow a reasonable
		definition of the mineralogy for the whole deposit. Following a mineralogical
		typing of the deposit, the expected properties of the product should be tested
		in order to determine which end uses and markets may be most usefully served.
		Iron stained or silica contaminated material should be subjected to
		metallurgical testing to determine the possibility of upgrading material at
		reasonable cost to increase the available product for high-end markets.





59



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Concurrently with the test work looking at
		physical and chemical properties, a market study should be undertaken to
		determine what nearby markets are available. A fuller market study would
		require pilot scale development of product so that potential consumers can be
		given material for testing in their particular application(s).

An advantage of this type of mineral
		operation is that once the deposit is defined and the markets reasonable
		assured, the production plans can be rapidly and cheaply advanced. We therefore
		anticipate that the suggested program could quickly be subjected to a nominal
		cost feasibility study. If this is positive production (qualified by the
		necessary permitting) can quickly be developed.



Drilling and sampling
			 
			 
			 
			 $
			 75,000Mapping and data compilation
			 
			 
			 
			 
			 10,000Metallurgy/Mineralogy
			 
			 
			 
			 
			 25,000Preliminary market
				  survey/sales
			 
			 
			 
			 
			 20,000Feasibility Study
			 
			 
			 
			 
			 30,000SUB TOTAL
			 
			 $
			 160,000Contingency 20%
			 
			 
			 30,000TOTAL
			 
			 $
			 190,000






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EXPLORATION TARGETS
		(OTHER THAN THE BURGIN)

At the current time there are only two
		mineral known zones which have the combination of grade and tons to have the
		potential to materially change the potential economics of the Burgin Mine
		development. These are known as Zone A and Ball Park.(See Figure No.7)Each of these mineral zones consist of a number of
		discrete but as yet poorly delineated mineralized bodies.

ZONE
		A

This area of lead/zinc/silver mineralization
		was discovered by means of an exploration drift on the 1050 level of the Burgin
		Mine. The purpose of the drift was to explore along the East Tintic Thrust
		fault to the north of the Burgin mineral deposits where it might intersect
		other northeast trending structures. The initial discovery was followed by
		approximately 15,000 ft of diamond drilling which outlined three areas of
		mineralization; respectively the 50-02, B-176 and Footwall.(These areas are shown in Figure No.7)A drive from the Burgin Mine, on the 1300 level, was
		designed to continue exploration of the A Zone. An underground
		course flowing 1,800 gpm of 140°F water was intercepted at the edge of the
		A Zone. All development on the 1300 level was suspended at the end
		of July 1977.





61



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62



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Martineau and Potter (1977) provide the most
		critical and detailed reportage of the Zone A mineralization. The
		description presented herein is based in large part on their report. A further
		report dated 1988 and written by, or for, Sunshine Mining Company, the lessee
		of the property, lists drill intersections for a number of the drill holes.
		These drill intersections and others are reported in the table below. None of
		these data have been verified by the authors of the current study.

SELECTED DRILL HOLE INTERSECTIONS FROM
		ZONE A

TABLE NO.12



Hole Number
			 
			 Intercept
			 
			 Pb%
			 
			 Zn%
			 
			 Ag oz/tonB-146
			 
			 151 - 167
			 
			 3.4
			 
			 9.4
			 
			 0.2B-147
			 
			 184 - 202
			 
			 6.2
			 
			 16.4
			 
			 0.2B-152
			 
			 177 - 95
			 
			 7.7
			 
			 5.0
			 
			 2.3201 - 221
			 
			 2.8
			 
			 0.8
			 
			 1.2221 - 242
			 
			 3.4
			 
			 1.5
			 
			 1.3242 - 283
			 
			 14.6
			 
			 4.8
			 
			 2.3304 - 313
			 
			 8.0
			 
			 0.4
			 
			 3.0B-161
			 
			 450 - 499
			 
			 4.2
			 
			 5.1
			 
			 2.7530 - 553
			 
			 5.7
			 
			 0.3
			 
			 2.8B-163
			 
			 339 - 405
			 
			 3.1
			 
			 6.8
			 
			 1.1B-167
			 
			 90 - 145
			 
			 1.6
			 
			 3.6
			 
			 Tr.145 - 190
			 
			 3.7
			 
			 4.7
			 
			 0.4278 - 360
			 
			 3.3
			 
			 3.2
			 
			 0.3B-176
			 
			 277 - 302
			 
			 23.1
			 
			 3.9
			 
			 4.1312 - 327
			 
			 9.2
			 
			 1.1
			 
			 3.6B-144*
			 
			 200 -
				  210(**)
			 
			 1.6
			 
			 4.2
			 
			 0.4B-150*
			 
			 170 -
				  195(**)
			 
			 4.3
			 
			 2.7
			 
			 0.6635 -
				  655(**)
			 
			 6.7
			 
			 11.7
			 
			 2.5745 -
				  765(**)
			 
			 9.7
			 
			 Tr.
			 
			 3.9B-151*
			 
			 340 -
				  350(**)
			 
			 1.3
			 
			 5.6
			 
			 0.3400 -
				  417(**)
			 
			 12.9
			 
			 22.2
			 
			 3.3B-168*
			 
			 320 - 70
				  (**)
			 
			 5.2
			 
			 10.6
			 
			 0.5B-169*
			 
			 095 -
				  145(**)
			 
			 4
			 
			 7
			 
			 Tr.180 -
				  230(**)
			 
			 1.8
			 
			 6.8
			 
			 Tr.360 -
				  385(**)
			 
			 5
			 
			 6
			 
			 3


*
			 Drill intersections and average
				  grades are taken from illustrations in Martineau and Potter (1977).


**
			 Intersections were measured on
				  figures from the Martineau and Potter (1977) report and should be expected to
				  have a low order of absolute precision.






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The 50-02 mineralization is located at the
		intersection of the north-east trending South fault zone and the East Tintic
		thrust fault. As such, it is similar to other important areas of mineralization
		such as the East Burgin, West Burgin and Tintic Standard deposits. The 50-02
		mineralization is somewhat unique in that extensive but erratic low grade
		lead/zinc mineralization containing minor silver credits was found by drilling
		in the basal volcanics and in the rubble zone below the volcanics. Similar
		grade, replacement style mineralization was found in the adjacent Ophir
		formation.

The grade of the mineralization is known
		only from company reports describing and evaluating the mineralization in the
		zone. The authors of this report did not find original assay certificates nor
		copies of the certificates. It is assumed the drift samples and core samples
		were assayed at the mine by Bear Creek staff, but this cannot be conclusively
		demonstrated. It must also be assumed that mineralization above the 1260 level,
		which was taken as a proxy for the water table will be partially oxidized and
		that metallurgical recoveries to be expected by froth flotation will fall
		somewhat short of those that might be achieved from milling a clean sulfide
		ore. No record of metallurgical testing of this mineralization was
		found.

Martineau and Potter (1977) assigned a
		reserve figure of 200,000 short tons to the 50-02 mineralization located
		between the 1050 level and the 1260 level. The estimated reserves were based on
		extending the values intersected on the 1050 level to depth and on the basis of
		a few diamond drill holes. They estimate the grade of the mineralization to be
		0.2oz/t silver, 4% lead and 6% zinc. They do not state what cut-off grade or
		other parameters were used in making this estimate. Elsewhere they noted that
		extensive additional drifting and drilling will be required to establish the
		tonnage and grade of any ore-body.

Martineau and Potter (1977) point out the
		potential to find additional, possible higher grade mineralization, with
		additional exploration in the 50-02 area both above and below the 1260 level.
		As an example of such potential they provide the example of the Tintic Standard
		Mine located approximately 700 metres (2300 feet) to the southwest. At that
		mine in the same fault zone and at a similar structural intersection 2.4
		million tons of ore were produced that averaged 24 oz/t silver and 12% lead.
		Grade and tonnage figures reported elsewhere for the same ore-body, as for
		instance by Shepard et al, (1967) are different, but of the same order of
		magnitude.

The B-176 area mineralization is located at
		the intersection of a north to northwest trending and west dipping reverse
		fault which displaces the East Tintic Thrust fault. The Tintic quartzite is
		brought into contact with the rubble zone at the base of the volcanics by this
		fault. Martineau and Potter comment that dolomites may underlie the quartzite
		below the 1260 level.

The geographic distribution and grade
		parameters of the B-176 zone are uncertain. The mineralization appears to be
		confined to a series of steeply dipping lenses which may merge with the 50-02
		mineralization. Similar depositions may exist in quantity below the relatively
		untested 1260 level.





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Martineau and Potter (1977) estimated the
		reserve tonnage of the B-176 mineralization located above the 1260 level to be
		150,000 tons at a grade of 3 oz/t silver, 10% lead and 5% zinc. They do not
		state what cut-off grade or other parameters were used in making this estimate.

The Footwall area mineralization is known
		from intersections of ten drill holes spread over an extensive area. The area
		is a very extensive mass of low grade lead/zinc/silver mineralization within
		the Footwall dolomites below the Tintic thrust fault. Within the low grade mass
		are pod like lenses of much higher grade mineralization for which lateral
		continuity remains to be established. The intercepts are interpreted as
		indicating a west dipping mineralization mass similar in character to the 274
		zone of the Burgin Mine.

Martineau and Potter reported a reserve for
		the Footwall mineralization located above the 1260 level. They estimated 60,000
		tons at a grade of 0.2 oz/t silver, 3% lead and 10% zinc. In addition they
		estimated a reserve of 240,000 tons located below the 1260 level. This reserve
		is estimated to contain 1 oz/t silver, 5% lead and 9% zinc. Again no cut-off
		grades or other limiting parameters are cited. They comment that if continuity
		of mineralization is assumed, the higher grade pods could contain up to 4
		million tons of the same grade material. They do not define
		higher-grade. At the same time they make the point that the
		potentially available tonnage of mineralization is quite large and at low
		cut-off grades (that are not otherwise defined) the Footwall area may contain
		more than 8 million tons with an estimated grade of 0.6 oz/t silver, 3% lead
		and 6% zinc.

The availability of large tonnages of
		lead/zinc/silver mineralization could provide some potential for lowering unit
		operating costs of production at the Burgin Mine if said tonnage allowed for a
		several fold scaling up of the rate of production and a longer period of
		production. It is likely that most of the Zone A mineralization is
		uneconomic if viewed in terms of current projected Burgin mining costs and
		smelter schedules. However, the grade, tonnage, metallurgical and mining
		parameters of this zone remain largely unknown. There is considerable potential
		for finding higher-grade mineralization and, following the 176 Zone
		mineralization down dip, might encounter higher silver grades. Martineau and
		Potter recommended and this writer concurs that an additional 7,500 ft of
		diamond drilling from the 1050 level is warranted to further define this zone.
		This drilling would, in addition to sampling for mineral grade, provide
		material from below the water table for metallurgical studies.

Recommendation

A two-phased exploration program is
		justified and recommended as a corollary to recommended development of the
		Burgin extension. A Phase 1 program estimated to cost $710,000 consisting
		largely of core drilling is proposed. The first exploration phase is to consist
		of 11,000 ft of diamond drilling from the 1050 level drift. A total of 7,500 ft
		will be used to trace the 176 zone mineralization down dip and 3,500 ft is to
		be used to follow up on Footwall mineralization high-grade intercepts. This
		work, which could follow the Burgin extension drilling, subject to its success,
		should establish the dimensions, grade and geological controls of
		mineralization in the 176 area and will





65



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continue to delineate higher-grade zones in
		the Footwall. The 176 area drilling should be sited to:

1.
			 Drill four holes spaced at 100 metre
				  intervals to intersect the 176 Zone at the 1360 level and


2.
			 Drill four holes spaced at 100 metre
				  intervals to intersect the 176 Zone at the 1700 level.


A preliminary metallurgical characterization
		study is included in the Phase 1 budget. The data acquired in Phase I should be
		sufficient for an initial engineering Scoping Study or included in the
		anticipated Burgin Feasibility Study. Provision is made in the Phase I budget
		for an initial engineering scoping study to determine the potential for
		mineralization identified in Phase I drilling to meet minimum economic
		parameters for production as either a stand-alone operation or as an adjunct to
		new production from the Burgin Mine.

The second phase of exploration in the 176
		Zone and Footwall should be dependent on a positive finding by the initial
		engineering study recommended for Phase I. Phase II exploration program should
		consist of 22,500 ft of drilling. Of this total 19,000 ft of drilling is
		recommended to be used to close the drill hole intercept spacing in the 176
		Zone to approximately 150 ft and to provide additional metallurgical sample
		material. A further 3,500 ft of core drilling is to be available for following
		up other potentially economic intersections or testing newly derived geological
		concepts.

Chief should also include more detailed
		metallurgical testing to supplement the Phase I testing. It is expected that it
		will be possible to produce a N1 43  101 Resource estimate based on the
		drill results generated from both Phase I and Phase II.

The budget for Phase II, at $1,312,000, also
		includes an allocation for additional metallurgical studies.

Cost Estimates (Excluding
		Metallurgy)



PHASE I - ZONE
				  A
			 
			 COST (US$)Diamond Drilling 3,500m @
				  $110/m
			 
			 
			 
			 $
			 385,000Assays and Analyses: 2,500 sample
				  @ $24
			 
			 
			 
			 
			 60,000Metallurgical Testing
			 
			 
			 
			 
			 25,000Drill Supervision and Sampling
				  (Geo, Living Exp., Transport)
			 
			 
			 
			 
			 90,000Report
			 
			 
			 
			 
			 35,000Engineering Scoping
				  Study
			 
			 
			 
			 
			 50,000Subtotal
			 
			 
			 
			 $
			 645,000Contingencies 10%
			 
			 
			 65,000TOTAL PHASE I
			 
			 $
			 710,000






66



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PHASE II - ZONE
				  A
			 
			 COST (US$)Diamond Drilling: 22,500 @
				  $35/FT
			 
			 
			 
			 $
			 790,000Assays and Analyses: 5,500
				  samples @ $24
			 
			 
			 
			 
			 132,000Metallurgical Testwork
			 
			 
			 
			 
			 100,000Drill Supervision and Sampling
				  (Geo, Living Exp., transport)
			 
			 
			 
			 
			 120,000Report
			 
			 
			 
			 
			 50,000Subtotal$
			 1,192,200Contingencies 10%
			 
			 
			 120,000TOTAL PHASE II
			 
			 $
			 1,312,000


BALL PARK
		ZONE

The Ball Park Zone of lead/zinc
		mineralization is located approximately 250 metres to the northeast of the
		northern edge of Zone A.(See Figure No.7)Faddies (1982) reports that the Ball Park
		mineralization was discovered as a result of a recommendation by D.C. Bulmer
		(Bear Creek Mining Company employee) after a program of surface mapping,
		geochemical studies and re-logging of 5000 feet of drill core.

Eight drill holes penetrated the
		mineralization and other drill holes intersected permissive and amenable
		structures or lithologies. The permissive dolomite is truncated to the west and
		northwest by the East Tintic thrust fault and to the south and east by a deep
		erosional valley now filled with Tertiary latite flows. The drill holes suggest
		the mineralization occurs in an area 600 feet wide by 1600 feet long and at a
		depth of at least 900 feet to 1000 feet below the surface. Faddies reports the
		deposit strikes to the north and dips to the east at 25° to 30°.(See Figures No.8 and No.9)





67



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68



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69



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The mineralization is believed to be a
		Mississippi Valley style lead/zinc deposit hosted by paleo-karst features in
		the Victoria Formation of Devonian age. This style of mineralization requires
		very close spaced sampling in order to confirm lateral continuity. In this case
		there appear to be three mineralized horizons in the dolomite. These are 35
		feet to 75 feet thick and are separated by 10 feet to 80 feet of barren
		dolomite. Faddies (1982) reports the better intercepts are below the water
		table.

DRILL HOLE INTERCEPTS OF BALL PARK
		MINERALIZATION

TABLE NO.13



Hole Number
			 
			 Avg. Thickness
			 
			 Lead percent
			 
			 Zinc percent
			 
			 Silver oz/tonET-971
			 
			 37.0
			 
			 1.76
			 
			 2.52
			 
			 0.1014.0
			 
			 0.41
			 
			 .047
			 
			 0.1012.0
			 
			 0.84
			 
			 2.10
			 
			 0.2035.0
			 
			 1.09
			 
			 3.34
			 
			 0.31ET-1051
			 
			 49.4
			 
			 0.20
			 
			 2.37
			 
			 0.1016.0
			 
			 0.23
			 
			 1.70
			 
			 Tr.12.0
			 
			 0.0
			 
			 3.30
			 
			 0.0114.0
			 
			 0.0
			 
			 1.04
			 
			 0.0122.0
			 
			 0.46
			 
			 3.45
			 
			 0.0016.0
			 
			 2.80
			 
			 6.80
			 
			 Tr.ET-1071
			 
			 28.0
			 
			 1.20
			 
			 4.03
			 
			 0.3942.0
			 
			 1.26
			 
			 5,02
			 
			 0.1359.0
			 
			 1.49
			 
			 4.15
			 
			 0.5426.0
			 
			 1.51
			 
			 41.6
			 
			 0.1126.0
			 
			 1.08
			 
			 5.38
			 
			 0.19ET-1202
			 
			 26.0
			 
			 0.52
			 
			 2.93
			 
			 0.0475.0
			 
			 1.12
			 
			 2.85
			 
			 0.2ET-1221
			 
			 32.02
			 
			 2.34
			 
			 6.68
			 
			 0.0037.02
			 
			 0.95
			 
			 4.19
			 
			 0.0010.02
			 
			 2.52
			 
			 8.90
			 
			 0.0026.0
			 
			 0.52
			 
			 2.93
			 
			 0.0475.0
			 
			 1.12
			 
			 2.85
			 
			 0.20ET-1671 ́
				  2
			 
			 23.0
			 
			 1.50
			 
			 3.90
			 
			 0.2010.0
			 
			 1.40
			 
			 5.20
			 
			 0.20ET-1682
			 
			 50.5
			 
			 0.6
			 
			 1.5
			 
			 Tr.119.0
			 
			 1.2
			 
			 2.4
			 
			 0.4ET-1692
			 
			 18.0
			 
			 1.5
			 
			 4.8
			 
			 Tr.ET-1702
			 
			 No mineralization to bottom of
				  hole, but may exist at depth.


1
			 Data as reported in Sunshine Mining
				  Company (1988)


2
			 Data as reported by Faddies
				  (1982)






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Faddies (1982) comments the holes completed
		in the Ball Park target indicate a potential for 13 million tons of
		mineralization with an estimated grade of 0.2 oz/ton silver, 1.3% lead and 2.7%
		zinc.

Sunshine Mining Company in its Special
		Report of 1988 cites a drill indicated mineral inventory Reserve at several
		lead+zinc combined cut-off grades.

BALL PARK ESTIMATES

(Varying Cut-off)

TABLE NO.14



Combined
				  Zinc+Lead1
			 
			 Average Grade
			 
			 Tons1
			 
			 Tons3Cut-off
			 
			 Zinc+Lead2
			 
			 Optimistic Estimate
			 
			 Conservative Estimate12%
			 
			 16%
			 
			 1,000,000
			 
			 200,0008%
			 
			 13%
			 
			 3,000,000
			 
			 700,0005%
			 
			 7.5%
			 
			 14,000,000
			 
			 3,000,0003%
			 
			 6.5%
			 
			 22,000,000
			 
			 4,900,000


1
			 Data from Sunshine Mining report of
				  1988


2
			 Estimates read from graph in
				  Martineau and Potter (1977)


3
			 Estimates read from graph in
				  Martineau and Potter (1977)


No information or justification is available
		as to the data used in the estimates. A comparison between Sunshine and
		Martineau and Potter (1977) shows a close similarity considering several
		additional mineralized holes were available to Sunshine in 1988.

The mineralization identified in the Ball
		Park deposit is relatively low grade and contains only negligible silver. There
		appears to be potential, with further close spaced drilling to identify
		substantial tonnages with a combined lead+zinc tenor in excess of 10%. Nothing
		is known of the metallurgical character of this mineralization, but good
		metallurgical recoveries are associated with Mississippi Valley deposits
		elsewhere. One might expect partial oxidation of sphalerite above the water
		table to cause poor recovery of contained zinc. Sphalerite recoveries from
		mineralization mined below the water table should be considerably higher than
		Kennecotts average zinc recovery from the Burgin Mine. Commonly,
		sphalerite from Mississippi Valley style deposits is low in iron and sells for
		a premium.

An engineering scoping study to determine
		potential viability of such a resource is warranted and should be carried out
		prior to any further drilling. A mining method designed specifically for this
		style of mineralization may prove to have substantially lower cost than those
		that pertain to the Burgin deposit. Room and pillar mining applied to these
		types of deposits in other districts have been able to achieve mining costs of
		less than US$7.00 per ton. The addition of several million tons of even
		marginally profitable ore delineated in the Ball Park could change the
		economics of the Burgin Mine development by reducing unit costs of
		production





71



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Recommendation

The company should commission an engineering
		Scoping Study, with an estimated cost of $50,000, to examine lower cost mining
		methods that might be applicable to Ball Park style mineralization and to
		determine an initial range of possible mining and milling costs for this zone
		of mineralization.

In the event that the range of costs and
		known metal grades combine to be suggestive of a reasonable potential of
		achieving positive economics by developing the Ball Park lead+zinc deposits,
		the company should develop a plan to gain additional knowledge as to the
		mineralization at this zone.





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OTHER EXPLORATION
		TARGETS

A large number of potential exploration
		targets can be identified with only a cursory review of maps of the Tintic and
		East Tintic districts. There are a number of well mapped faults, the
		intersections of which might be expected to produce the requisite porosity and
		permeability to permit emplacement of mineralization, given reactive rocks and
		the presence of ore forming fluids. The recognition of the permissive
		environment at the intersection of high angle faults and the thrust faults
		provides a key target area that has been successfully exploited in the East
		Tintic district. Exploration in Zone A identified, for perhaps the first time
		in the district, large amounts of low grade mineralization where the porosity
		and permeability of the rubble zone and paleosoil at the base of the Packard
		Quartz Latite are intersected by the northeast trending hydrothermal fluid
		bearing fissures.

Notwithstanding the ease of recognizing
		structural intersections, a number of major and junior companies have explored
		in the district with little success. Kennecott Minerals Company and Bear Creek
		Mining Company (a Kennecott subsidiary) in contrast demonstrated marked success
		through application of geological principals to develop, qualify and prioritize
		targets of merit. This same process needs to be applied to the many known
		structural intersections in order to precisely locate the most permissive
		environments for mineralization and the most permissive environments for
		mining.

There is at present, based on a four day
		examination of the data base at the Burgin Mine, a paucity of well documented
		and qualified targets that warrant immediate exploration expenditures. A number
		of structural intersections have been mentioned in reports, but detailed
		descriptions allowing prioritization of drill or exploration targets is
		generally lacking. Nowhere is there a detailed description of the targets such
		as the expected depth to the permissive horizon, the permissive unit that is
		the specific target of the exploration, the attitude of the enabling structure,
		the degree or type of alteration and how it compares with other ore zones in
		the region or with the alteration to be expected in the productive stage
		alteration facies, or a series of arguments and observations that could be used
		to qualify and prioritize potential targets. This situation is not a
		consequence of a lack of exploration potential, but rather a result of the
		focus of time and effort in recent decades on the development of the Burgin
		Mine and deposits in its immediate vicinity.

We recommend the application of modern day
		technology in the structural analysis field to the Tintic Districts. A
		concerted effort should be made to digitize all the available data accumulated
		over the years. This data can then be represented three dimensionally to
		produce exploration targets and to opine as to their potential size and
		priority. This aspect for the future should be considered seriously.

Recognition of regions with exploration
		potential in the Tintic District is a relatively simple and intuitive process
		for an experienced exploration geologist. For example, virtually any
		intersection of two faults is a potential target. Over the years the following
		targets have been identified:





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1.
			 A.C.A. Howe selected as worthy of
				  comment, the following exploration targets:



			 756 Fissure narrow gold-silver
				  target in the Trixie Mine area.



			 20th
				  Century-Trixie-Eureka Standard Fault Intersection.



			 Eureka Standard mine below water
				  level target where mining of 2 ft to 6 ft wide veins was stopped by water
				  incursion and potential low grade target of sub-10 g/t cut-off mineralization
				  left in previously mined levels.



			 South Apex Hill Drill Hole ET-126
				  area of silicification and pyritic alteration and low-grade
				  mineralization.



			 Iron Blossom large vertical pipe of
				  86 g/t silver mineralization left in underground workings and



			 Zuma Area gold target in a zone of
				  halloysite clay alteration.


2.
			 R.E. Irwin (2001) selected another
				  suite of exploration targets:



			 Inez fault and East Tintic thrust
				  fault intersection target for Burgin type lead/zinc replacement
				  deposits.



			 50-02 Area lead/zinc replacement
				  deposits (discussed in detail as part of Zone A above).



			 Tintic Standard offset where the
				  Tintic Standard thrust fault is offset along the South fault producing a
				  lead/zinc replacement target.



			 Homansville target (only partially
				  on the Companys claims) for steeply dipping fissure/vein
				  gold/silver/copper deposits similar to the Trixie Mine.



			 Baltimore and North Lily Fissure for
				  steeply dipping fissure/vein gold/silver/copper deposits similar to the Trixie
				  Mine.



			 Iron King Area of silicification and
				  iron-manganese staining in the hanging wall of the Eueka Lily fault.



			 Zuma Clay Pit Area gold
				  target.



			 East Tintic Consolidated Area
				  intersection of the Eureka standard Fault and the 20th Century fault
				  vein hosted gold/silver target.



			 Trixie West Target vein hosted
				  gold/silver target.



			 274 Zone lead/zinc replacement east
				  of the Burgin Mine in the footwall of the East Tintic thrust fault.



			 Zone A lead/zinc replacement.



			 Central Standard/Homansville (not on
				  the Companys claims) five drill holes are reported to carry narrow
				  intersections of low grade silver/lead/zinc mineralization.



			 Oxen-Tiptop Manganese Occurrences
				  are believed by Irwin to be an alteration halo of a concealed base metal
				  deposit.



			 Ball Park Mississippi Valley
				  lead/zinc target.


In 1988 Sunshine Mining Company geologists,
		Messrs Glen Mellor and Ralph Stitzer and Engineer Allan Young, after several
		years of exploration work on the property identified the following resource
		potential:





74



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DISTRICT TARGETS AND POTENTIAL

TABLE NO.15



NAME
			 
			 PROPERTY
			 
			 RESOURCE POTENTIAL1. New Burgin Ore-body
			 
			 Burgin Lease
			 
			 1,284,406 tons @ 22.57 OPT Ag, 19.5%
				  Pb, 5.53% Zn2. Burgin Gold
			 
			 Burgin Lease
			 
			 500,000 tons @ 0.25 OPT Au, 10 OPT
				  Ag3. Burgin Tailings
			 
			 Burgin Lease
			 
			 1,168,655 tons @ 1.3 OPT Ag, 2.45%
				  Pb, 2.75% Zn4. Ball Park
			 
			 Unit Lease
			 
			 30,000,000 tons @ 1.0 OPT Ag, 3-4%
				  Pb, 4-6% Zn5. Zone A
			 
			 Unit Lease
			 
			 450,000 tons @ 1.5 OPT Ag, 6.4% Pb,
				  7.8% Zn6. 50-02 Area
			 
			 Unit Lease
			 
			 200,000 tons 0.2 OPT Ag, 4% Pb, 65
				  Zn7. 756 Fissure
			 
			 Unit Lease
			 
			 200,000 tons @ 0.25 OPT Au, 8 OPT
				  Ag8. Inez Area
			 
			 Unit Lease
			 
			 500,000 tons @ 0.3 OPT Au, 5-8 OPT
				  Ag9. South Apex Hill
			 
			 Unit Lease
			 
			 100,000 tons @ 0.25 OPT Au, 5-8 OPT
				  Ag10. 20th Century-Trixie
				  Int.
			 
			 Unit lease
			 
			 150,000 tons @ 0.25 OPT Au, 5-8 OPT
				  Ag11. Eureka Standard Deep
			 
			 Unit Lease
			 
			 300,000 tons @ 0.8 OPT Au, 12 OPT
				  Ag12. Eureka Standard Mine
			 
			 Unit Lease
			 
			 100,000 tons @ 0.2 OPT Au, 3-5 OPT
				  Ag13. Eureka Standard Fault
			 
			 Unit LeaseBurgin Lease
			 
			 200,000 tons @ 0.2 Au, 7-10 OPT
				  Ag14. Middle Fault
			 
			 Burgin Lease
			 
			 100,000 tons @ 0.2 OPT Au, 7-10 OPT
				  Ag15. Iron Blossom Pipe
			 
			 Iron Blossom Lease
			 
			 150,000 tons @ 0.2 OPT Au, 3-4 OPT
				  Ag16. Colorado No.1
			 
			 Iron Blossom Lease
			 
			 65,000 tons @ 0.15 OPT Au, 4-7 OPT
				  Ag17. Pine Canyon
			 
			 Iron Blossom Lease
			 
			 100,000 tons @ 0.15 OPT Au, 4-7 OPT
				  Ag18. Inez-Thrust Int.
			 
			 Burgin Lease
			 
			 1,000,000 tons @ 15 OPT Ag, 15% Pb,
				  5% Zn19. Zuma Area
			 
			 Zuma
			 
			 300,000 tons @ 0.3 OPT Au, 8 OPT
				  AgUnit Lease20. Tintic Standard Offset
			 
			 Unit Lease
			 
			 1,000,000 tons @ 20 OPT Ag, 20%
				  Pb21. Iron King Area
			 
			 Unit Lease
			 
			 250,000 tons @ 10 OPT Ag, 10% Pb, 5%
				  Zn22. Waste Dumps
			 
			 Iron Blossom LeaseUnit LeaseBurgin Lease
			 
			 255,000 tons @ 0.025 OPT Au, 1.56
				  OPT Ag23. Trixie Mine
			 
			 Unit Lease
			 
			 250,000 tons @ 0.2 Au, 5-6 OPT Ag






75



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NEAR SURFACE
		EXPLORATION POTENTIAL

District in
		General

During the period 1986  1987 Western
		Mining Company (WMC) of Australia spent time in the Tintic District. Allegedly
		they left, not because of any lack of interest, but because of difficulties in
		partnership relationships. Five well known and respected geologists spent time
		in the district and Ian W. Levy authored a report of their findings.

Specifically, relative to gold and near
		surface targets the report stated,Exploration in the past has focused on silver
		 lead and gold ores rather than gold specifically. Attention has centred
		on highest ore value areas, which are generally silver  rich base metal
		ores. Consequently, the areas most favourable for gold exploration are
		under-explored and an intensive exploration project is recommended. The project
		should use empirical exploration methods (geochemistry and geophysics) plus
		conceptual exploration methods (e.g. structural and lithological
		modeling).

The following comments, possibly of
		importance for future development in the district, especially in regard to open
		pit targets are extracted from the WMC report.


			 Even though more than 2.7 million
				  ounces of gold has been produced, gold is rarely visible in both ore and stream
				  sediments because it is fine grained and associated with sulphides and silica
				  (see Tower and Smith, 1899 p. 685). Post prospecting could have missed surface
				  silica  gold deposits .



			 Analogous mining districts are El
				  Indio in Chile, Lepanto in the Phillipines and the Nansatsu  Ishihari
				  gold belt in Japan.



			 In the main Tintic District, gold
				  grades increase to the south. This pattern does not hold true for the East
				  Tintic District.



			 Gold grades increase from east to
				  west.



			 When gold and tons are combined
				  two zones are highlighted, which suggests two mineralizing centres or intrusive
				  stocks.


1.
			 Centennial  Eureka to
				  mammoth area.


2.
			 Eureka Standard to Tintic Bullion
				  and North Lily.



			 WMC felt that from 1955 emphasis had
				  been placed on the East Tintic District and gold rich targets had not been
				  explored in any detail.


The concluding section of the WMC report is
		repeated below.





76



---



Exploration and
		Prospectivity

Whilst the Tintic District was being
		extensively mined, exploration was mainly done by underground development and
		known ore runs were often pursued to the exclusion of other prospective areas.
		However, in areas covered by mine development, exploration was rigorous and the
		miners knew of the indictor mineral leads which could be followed to ore zones.
		Mining generally continued through to barren limestone, especially the
		scavenging of siliceous ores by lessee miners after 1927.

Major mining companies including
		Anaconda, Kennecott, Asarco and Exxon have explored in the district but all
		except Exxon targeted their efforts towards silver  rich lodes. The Exxon
		gold exploration at Treasure Hill was part of a porphyry-copper and related
		deposits exploration project. Virtually none of the exploration done since 1935
		has been for gold  the district zoning patterns presented in Section 6.3
		explain why very little exploration has been done in the areas of major gold
		enrichment. Most of the areas considered prospective in this report were not
		part of the leases explored by these major companies.

The main gold potential lies in new ore
		zones, either away from or above old workings and in prospective areas not
		previously tested. Surface drilling of the prospective areas has been totally
		insignificant. The inherent difficulty of drilling altered limestone terrain is
		recognized as an exploration inhibitor (sic) in the past, but modern drilling
		mud technology and drill casing will permit efficient testing of new target
		areas.

Surface prospecting for gold has been
		ineffective because the gold is generally very fine and has rarely been seen in
		gold pan concentrates. Modern geological and geochemical techniques must be
		tried over this major mineral field. It appears that chip sampling of surface
		structures would be a very cost effective exploration method to employ.
		Anomalies thus identified should be diamond drilled to gain maximum geological
		knowledge and any encouragement should be followed-up by the appropriate
		percussion drilling. Resistivity surveys may help identify silicified zones.
		The potential for shallow concealed deposits is very high.

A vast number of Tintic mining tenements
		have been combined into a significant joint venture exploration area for the
		first time and data from almost all previous operators have been made available
		to officers of WMC, thus presenting a unique opportunity to explore more
		effectively.

Exploration
		Targets

Figure 23shows the areas considered most prospective for major
		gold deposits. Specific exploration targets identified from studies done with
		Craig Reddell are:

1.
			 Centennial Eureka south of the
				  shaft above the 8 level to the surface.






77



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2.
			 Above the auriferous
				  granite pipe zone of Chief Consolidated Mine on the Eagle Channel
				  of the Mammoth-Chief Ore Run. Geochemically anomalous vein/fault material has
				  been found beneath the water tanks behind the Chief offices and no mining
				  exists above the 600 foot level.(See
				  Figure 24)






78



---







Selection of specific drill sites is a
		somewhat more difficult task and has to be based on a





79



---











80



---



detailed understanding of complex but
		local geological relationships. Prioritizing of exploration targets in order to
		have the best chance of achieving an economic mineral discovery is perhaps even
		more difficult and requires a detailed knowledge and understanding of
		geological relationships across the district. Prioritizing exploration targets
		is a process that must be completed to maximize the potential for discovery and
		to make the most efficient use of fiscal resources.

The target represented by any particular
		structural intersection needs to be substantially refined to be a useful target
		for the following reasons. The sites for mineral deposition are at substantial
		depths of 400 metres or more below the surface. The target mineral zones are in
		most instances relatively small, irregular bodies that in many cases have only
		slightly larger alteration aureoles. The maximum dimensions of the targets are
		typically less than 250 metres and it is not uncommon for considerable ore to
		be produced from bodies that have cross-section al dimensions of less than 30
		metres. These bodies are not easy targets to find. Surface mapping by the
		United States Geological Survey and by others has identified alteration
		patterns in surficial rocks that in a few cases have been traced down structure
		to ore mineral zones at depth and a sequence of alteration patterns has been
		developed which allows recognition of alteration associated with mineral
		deposits. Geophysics seems to have been used but little in the district to
		qualify targets at depth. Recognition that specific stratigraphic units host
		the majority of the ore producing mines suggests that it is possible to select
		high probability targets by focusing on the intersection of the most permissive
		units with the fluid carrying fissures and/or the porosity generating
		faults.

Mining below the water table is
		substantially more complex than above the water table and hence the potential
		for quantities of ore above the water table should be important criteria for
		prioritizing exploration targets. The quality and quantity of water vary
		importantly from place to place in the district. Some of the acquifers in the
		district, such as that near the Burgin Mine, are thermal waters rich in
		dissolved salts. Disposal of such fluids becomes extremely costly. The cost of
		pumping water out of the mine openings and more recently regulatory issues with
		water discharge have caused the halting of mining in some mines and has
		effectively sterilized other discoveries. The elevation of the water table
		varies throughout the district, but is generally known from the many bore holes
		and mine openings.

A series of plans and sections should be
		generated using all available private and public information and which clearly
		show the geological relationships and information mentioned in the two
		paragraphs above. The data presentation should be in such a manner as to
		clearly delineate which information was observed, which was derived from
		interpretation or extrapolation from nearby observations and which information
		is mere speculation. The plans and sections should be used to identify specific
		exploration targets for drilling and to provide the information for an initial
		prioritization of drill targets. Where warranted, geophysics should be applied
		to further refine high priority targets to achieve final selection of drill
		collars and targets. It is likely to take two geologists between six months and
		a year to generate a rational synthesis of district geology and to select and
		prioritize the most prospective exploration targets.





81



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Trixie Ore

Several piles of as  mined and crushed
		Trixie ore are located on surface near the mine shaft and concentrator. The
		estimated tonnage and grade (based upon assays of the broken ore remaining
		underground) of this material is as follows:



120  150 tons @ 0.67 opt Au;
				  5.3 opt AgMINE400  500 tons @ 0.18 opt Au;
				  1.7 op6t AgMILL UNCRUSHED
			 
			 750 tons @ 0.67 opt Au; 5.3 opt
				  AgMILL CRUSHED
			 
			 120 tons @ 0.67 opt Au; 5.3 opt
				  Ag


The contained gold content is approximately
		750 ounces as well as 6,100 ounces Ag. Assuming a 90% gold recovery and 75%
		silver, with present day metal prices of $450 Au and $7.50 Ag this equates to
		(approximately) a concentrate value of $335,000.

Burgin
		Tailing

Approximately 1.2 million tons of tailing
		containing some 2.5% Pb; 2.8% Zn and 1.3 opt Ag have been estimated to remain
		on the property. With current metal prices of $0.70/lb for Zn, 0.45/lb Pb and
		$7.50 per ounce Ag, this equates to an in-situ gross value of $86MM. Much of
		this material will be difficult to recover to concentrate, non-the-less it
		remains a target for future metallurgical and process investigation.





82



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ENVIRONMENT/PERMITTING

This essential aspect of the on-going
		development at Chiefs Tintic property has not been investigated for this
		study. It is our understanding that Chief has arrived at an acceptable solution
		with the EPA regarding an environmental problem in the Main Tintic
		District.

We also understand that the State of Utah
		has granted three permits for operations in the District.


			 Notice of Intention for the Burgin
				  Mine, which was issued on June 3rd 1985 to Tintic Utah Metals LLC
				  (M/049/009).



			 Notice of Intention for the Trixie
				  Shaft Area which was issued December 10th 1993 to South Standard
				  mining Company (M/049024).



			 Notice of Intention for Trixie West
				  Exploration Permit Number E/049/046 was issued to Chief Gold Mines Inc. on July
				  16th 1995.


Any future program of work should include an
		updating of the benefits and obligations under these permits as well as a full
		definition of the environmental risks associated with any proposed operations.
		Equally as obvious is the need to understand and comply with all environmental
		and permitting requirements for the work proposed in this study.





83



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MAIN REFERENCE
		LIST



1.
			 Volumes of files at Chiefs
				  Eureka office and the Burgin office.2.
			 Mine Development Associates; Updated
				  Feasibility Study, October 5, 2001.3.
			 Morrison Knudsen; Review and Audit
				  of Sunshine Mining Companys New Burgin Project.4.
			 Pincock Allen and Holt Inc.;
				  Estimate of Mineable Reserves for the New Burgin Project, August 1,
				  1989.5.
			 Glenn M. Mellor; Resource Potential
				  for Sunshine Mining Companys East Tintic Properties, January 15,
				  1990.6.
			 Donald B. Tschabrun; Resource
				  Estimate for the Burgin Project, November 11, 1997.7.
			 Donald B. Tschabrun; Interim
				  Resource Estimate Burgin Project, July 30, 1996.8.
			 Various submittals to the Utah State
				  Department of Natural Resources for the Burgin Water Appropriation.9.
			 Martineau and Potter; Ore Potential
				  at the Tintic District, May, 1977.10.
			 Paul Spor; Chief Application to
				  Appropriate Burgin Mine Water, 2000.11.
			 Various Submittals to the State of
				  Utah Engineer, 1998  1999.12.
			 Chester Engineers; Draft Feasibility
				  Study Burgin Water Desalination, October, 1998.13.
			 Hal Morris; The Main Tintic Mining
				  District.14.
			 Sunshine Mining Company; Holdings in
				  the East Tintic District February 12, 1987.15.
			 H.T. Morris & T.S. Lovering;
				  General Geology and Mines of the East Tintic Mining District, 1979.16.
			 Bear Creek Mining Company;
				  Exploration Project Review  Tintic, April, 1982.17.
			 Western Mining Company; Shallow Gold
				  Potential Tintic Mining District, February, 1987.18.
			 Hal T. Morris, Paul Mogenson; Tintic
				  Mining District, 1978.






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