Search Documents

By O. Alonso-Gonzalez, F. Nava-Alonso, A. Uribe-Salas, R. Pérez-Garibay, C. Jimenez-Velasco

A current problem in gold production is the presence of copper minerals which dissolve during conventional cyanidation practice. The dissolved copper causes several problems in the different stages of the process. This issue is the base for copper elimination/recovery methods such as the Sulphidization, Acidification, Recycling, and Thickening (SART) process, which is currently used in gold production plants around the world. Even though the process is well established, new challenges appear as the composition of the solution varies from ore to ore, in some circumstances affecting the performance of the SART process. This work presents a thermodynamic study and experimental tests performed to understand the role of the initial copper concentration and sulfur addition, on the efficiency of the copper removal. For a constant cyanide concentration, an increase in the initial copper concentration causes a decrease in copper removal; high sulfur/copper molar ratios favor copper removal. We demonstrated that the cuprous sulfide formed from solutions with high ionic strength, presents better sedimentation rates than solutions with less ionic strength. For solutions with low sedimentation rates, the use of some pretreatment is necessary to facilitate the sedimentation such as addition of flocculants, particle recirculation and the used ultrasonic pretreatment.

Jan 1, 2014

By Gordon E. Agar

"The effectiveness of two this compounds for copper sulphide depression in molybdenite flotation has been examined with batch bench scale tests.Both thioglycollate and trithiocarbonate proved to be good depressants producing results similar to those f or cyanide and hydrosulphide. Ferrocyanide was ineffective on the samplematerial tested.The depressing effect of thioglycollate was readily reversed by an addition of xanthate.IntroductionThe separation of molybdenite from copper sulphide minerals has been practised for a long time(1,2). The predominant scheme is to depress the copper sulphides while capitalizing on the ready floatability of molybdenite to effect the separation. There are many recipes for depressing the copper minerals(3), among them the addition of sodium cyanide, particularly if the copper is present primarily as chalcopyrite; some form of sulphide whether as sodium sulphide, sodium hydrosulphide or phosphorous pentasulphide when there is a mixture of copper sulphides such as bornite, chalcocite and chalcopyrite; potassium ferro cyanide may be used when chalcocite is the principal copper sulphide; peroxide or hypochlorite; as well as heating in the presence of lime. There are different mechanisms by which these reagent s depress the copper sulphides. Sodium cyanide and the sulphide-producing compounds apparently reduce the potential of the copper sulphides to the point where they ar e rendered unfloatable by xanthates . Another study(4) showed that after cyanide addition, chalcopyrite was rendered un floatabl e until the REDOX potential (measured with a gold electrode with a saturated calomel reference) had become more positive than -250 mY. If the chalcopyrite pulp was conditioned after the cyanide addition the REDOX potential increased slowly, whereas if air was introduced during the conditioning the potential r ose much more rapidly. The difference in time to reach the desired level of - 250 mV was as much as 20 minutes' 25 minutes being required when air was absent but less than 5 minutes when air was admitted . These observation s demonstrate some of the problems encountered when using cyanide to depress sulphides such as chalcopyrite . The system adjusts to counteract the depressing effect of the additive. To overcome this tendency it is necessary to either add an overdose at the beginning or to more or less continuously add the regulating reagent. In this study cyanide solution was added under automatic control to maintain a specified REDOX potential."

Jan 1, 1984

By Robert W. Bartlett

Chalcopyrite in copper flotation concentrates can be hydrothermally converted to digenite with oxygen in a continuous process, with rapid kinetics when operated above 200°C. This eliminates large amounts of sulfur and iron. These "super-concentrates" have potential for debottlenecking smelters and avoiding capital intensive future smelter expansions, otherwise required in the late 1990s. The technology of producing super- concentrates is described, with cost estimates, and the use of super-concentrates as smelter feed is discussed, with examples.

Jan 1, 1992

By Alexander Sutulov

1. INTRODUCTION According to the classical economic approach of the open market economy, copper supply should be regulated by market forces which are (a) demand and (b) prices. In fact, demand is of primordial importance because prices are a reflection of supply-demand balance: they go up in times of shortages of supply and down when over-supply situation exists. In the past few years, this basic mechanism of market economy has been under pressure from a number of non-market forces, such as social and monetary problems of sane important copper producing countries; structural changes in copper mining companies through mergers with oil companies; and other factors which greatly contributed to distortion of the basic supply-¬demand equation. As a result, we have faced in the last three years a situation of oversupply rarely seen in the past, with a catastrophic fall in copper prices, huge losses at mines and unprecedented indebtedness of copper producers. This has deep repercussions in the copper business and its future supply outlook. This is particularly so when taking into consideration that high indebtedness and high interest rates conspire against the very survival of some traditional producers. As we write this comment (July of 1985), close to one half of U.S. copper production capacity is closed down, sane of it for good, while the same precarious situation exists in Canada, Australia, Zambia, and some other primary producers who cannot face present low prices with equally low costs. The question is then what this situation means for the medium-¬term future, until the end of this decade, Particularly when taking into consideration the fact that some of the factors influencing it are of a long-term character. 2. SUPPLY-DEMAND SITUATION IN THE LAST FIVE YEARS Since the last boon year, in 1979, when world copper consumption reached a record of 9.8 million metric tons of refined copper, demand dropped to only 9.0 million tons in 1982, when the crisis bottomed out, and then steadily recovered to the level of 9.7 million tons in 1984. Meanwhile, copper production, which in 1979 was only 9.4 million tons, in spite of the fall in demand, grew to 9.7 million tons in 1981 and 1983 and only temporarily was cut to 9.4 million tons in 1982 and to 9.5 million tons in 1984. As a result of this imbalance, hugh copper stocks were accumulated with supply systematically out¬growing demand by over 2 million tons at one time. With the Western World refined copper demand running at an average of 115,000 tons per week, this oversupply was equivalent to between 15 and 18 weeks production, i.e. twice as large as normal. This was immediately reflected in very low copper prices. Another contributing factor to relatively low copper prices was the lack of speculator interest for this commodity. Between 1981 and 1984 the world experienced a very intensive period of wealth distribution, high interest rates and intensive restructuring off world economy and its industrial sector. Therefore, much of the available money was attracted to more lucrative and attractive investments rather than to the lack-luster performance of copper, and for that matter, of the metal sector in general. The much announced and discussed recovery of the U.S. economy since 1983, and particularly in 1984, with an overall growth of the GNP by 6.8% in 1984, happened to be more oriented to the service sector than the industrial sector and therefore this recovery was not adequately reflected in copper consumption nor to metals in general. This put serious doubts about the future demand for copper. Consequently copper prices failed to improve in spite of an apparent fall in stocks and a healthy increase in the demand for copper since 1982. At the time of this writing (July, 1985), although same analysts claim that the available stocks are running at about 8 weeks of supply, which are considered to be below normal, copper prices are still very low at between 65 and 67 cents per pound of copper. Part of the answer may be the normal seasonal slow-down period in the Northern Hemisphere due to summer vacations and

Jan 1, 1986

By T. Mutoh

Tellurium in raw slime has recently been recovered as copper telluride in Onahama Smelter & Refinery. The slime is decopperized with sulfuric acid solution in conventional leaching tanks. The pregnant leach solution containing 45 g/l of copper and 1 g/l of tellurium is introduced to a copper telluride recovering reactor. The tail solution from the reactor containing less than 20 mg/l of tellurium is recycled directly to circulation tanks of copper electrolysis. The impurities of the product copper telluride is kept below 1% without any purification.

Jan 1, 1990

By M. Smit, N. C. J. Dempers, S. Lunkov, C. Saguru, J. Nel

A common method for the recovery of copper is by heap leaching, solvent extraction, and electrowinning. During solvent extraction, the solution is continuously recycled, and it is necessary to remove impurities that build up in the raffinate. For the process investigated in this study, the main impurities include Fe, Co, As, and Zn. Recovery of these impurities can provide economic benefits while adhering to legal and environmental constraints. Such impurities are often removed by precipitation utilizing a neutralizing agent, such as limestone or lime, but it is also possible to remove the impurities by sulfide precipitation. Sodium hydrosulfide (NaHS) is used to selectively precipitate copper, arsenic, and zinc from a solvent-extraction raffinate stream. The resulting sulfide precipitate is bagged and sent to a roaster for further treatment. The design of the sulfide precipitation was investigated to optimize the processing conditions (including temperature, pressure, and residence time), the materials of construction, and the toxicity and explosive nature of the hydrogen sulfide gas produced

Jan 1, 2023

By T. Shellhammer, A. Park

Among all pure metals, copper has the second highest thermal and electrical conductivity, next to silver. These properties coupled with a reasonable cost, make copper ideal for use in metallurgical vessels as a conductor of electrical current, as well as a lining material and protector of refractories and steel superstructures. Treatments of copper to address it’s softness and low melting point, specifically hard surfacing and calorizing, will be discussed. This paper will cover the evolution of copper to its current form, offer comparisons to other materials, and theoretical and empirical evidence on its superior performance in metallurgical vessels such as blast furnaces, basic oxygen converters and electric arc furnaces. Some case studies will be provided and finally a brief summary.

Jan 1, 2015

By Clinton H. Crane

WHEN I agreed to accept your Committee's suggestion that I talk to you at this meeting on the future of copper, lead and zinc mining, I could not help wondering why I had been chosen to be a prophet. Knowing the pious nature of your committee, I thought I had better look up and see what the Bible says about being a prophet, and what the dictionary definition of a prophet is. The word "prophet" comes from the Greek and is defined as "one who speaks through a divine inspiration." Also, in the Gospel of Saint Matthew, Christ tells us that a prophet is "not without honor save in his own country and in his own house." That gave me pause until I remembered that after all, I had started my professional career as a Naval Architect, not as a Mining Engineer, and consequently, this might be a clue to the fact that the eminent geologists of this society, who know infinitely more about the world's crust than I do, were passed over in this assignment. In thinking back, I then remembered that the first really successful prophet was Noah and that Noah was a Naval Architect. Here then was the answer. One more quotation from the Bible I shall take as my text: In Ecclesiastes, chapter I, verse 9: "The thing that hath been is that which shall be." Now, with this preamble, and not being bound by anything save inspiration, I will proceed to prophesy: As far as the next 15 to 25 years are concerned, from the studies made by

Jan 1, 1947

By A. Block-Bolten

Copper concentrates from northern Peru contain large amounts of arsenic and antimony in the form of sulfosalts like enargite, tennantite and tetrahedrite. These impurities affect up to 30% of Peruvian copper production. Under the sponsorship of the Instituto Geologico Minero y Metalurgico in Lima a project to separate copper from arsenic without pollution of the environment is being carried out at the National University of Trujillo in Peru. The U.N.T. process is unique in the respect of using .low melting salts as solvents, and sulfides of metals with high cation potential as destructors of the complex sulfosalt molecules. Laboratory scale experiments have given encouraging results but interesting and difficult problems developed on scale-up. The process is pyro-hydrometal-lurgical and under certain circumstances its economics are promising.

Jan 1, 1981

By Dachun Liu, Baoqiang Xu, Bin Yang, Yong Deng, Heng Xiong, Xiumin Chen

"In this paper, vacuum distillation technology has been employed to study the possibility and regularity of separating copper and enriching lead, silver and other valuable metals from copper- based multi component alloys. The result of experiment indicated the practicability of vacuum distillation dealing with copper base complex multi-component alloy to separating lead and silver, when the vacuum distillation temperature was 1637K, distillation time was 60min; residual gas pressure was 20-30Pa, the removal rate of silver was 95.86%, of lead was 99.63%; the evaporation rate of copper was 0.745%. The removal rate of arsenic, antimony and tin respectively were 83.09%, 48.12% and 24.54%. Vacuum distillation method is applicable for dealing with the copper-based multi component alloys, silver and lead can realize effective recovery.IntroductionCopper generally go with lead mineral, therefore, to remove copper from lead becomes an inevitable link in lead smelting. Dross from removed copper usually contains lead, arsenic, antimony, tin and noble metal silver. In abroad, soda ash-iron dust is adopted to process copper dross of this kind. The outcomes of this process are lead matte, wet lead, scoria smoke dust etc[1~5] The dominant components of all the outcomes are copper and lead. Moreover, in the process of tin smelting, copper contained by-products are produced, such as, crude tin alloy, tin lead electrolytic anode slime and so on. Copper content of all the products are ranging from few percentages to several tens percentage. Those products, no matter merge into main producing process, or separate treatment cost losses of valuable metal. For one hand, the containing of copper of outcomes from pyrogenic attack is less than 50%, for the other hand, a lot of valuable metal need to be processed, which increase the cost of retrieve."

Jan 1, 2013

By Zay Jeffries

AN acute current shortage of copper, with the prospect that conditions may become worse, indicated by Office of Production Management information. Present estimates of copper requirement for defense indicate that nearly all the available metal will be required in the future for this purpose. Severe dislocations are, expected, but the objective is to distribute the copper available for civilian use in such a manner as to minimize these dislocations. The advisory committee on metals and minerals has been asked for a suggestions which may be helpful in meeting this situation.\

Jan 1, 1941

By W. H. (Bill) Jay

The destruction of cyanide creates a significant impost on the financial returns of gold mines. These losses increase substantially when ores containing cyanide-soluble copper are treated. Failure to remove or destroy free or WAD cyanide before it is discharged into tailings dams has, on a number of occasions, led to significant fish, bird and animal deaths. Thus, the retention of cyanide within the leach circuit and the recovery or removal of stable metal cyanide complexes needs to become standard practice for all gold mining operations. Furthermore, reducing cyanide purchases will minimise the quantity of cyanide transported to the mine site thereby reducing the opportunity for spillage of this chemical. Many different physical and chemical methods have been investigated for the removal of cyanide from process solutions. All cyanide destruction methods create a financial impost on the economic viability of the gold mine. To reduce these costs, the recycle of cyanide and the economic recovery of the metals solubilised by cyanide is of prime importance. The alternative methods currently in use, or under consideration for the treatment of copper cyanide, entail high reagent costs, produce the recovered metal in a contaminated form and require the safe handling of hydrogen cyanide. These processes will be briefly discussed within this report Environmental and financial benefits are enhanced by recent laboratory and pilot plant studies conducted by Oretek Limited. The Oretek CPC process for the recovery of copper and the recycle of cyanide from copper cyanide-containing solutions does not require pH adjustment to the alkaline cyanide leach conditions. Copper ions are coordinated to a water-soluble and/or an organic soluble polychelating agent under alkaline conditions and the cyanide ions are displaced Membrane separation, or phase separation of the polychelated copper from the cyanide ions enables the cyanide ions to be recycled and the copper to be directly electrowon from the polychelating polymer. The polychelating polymer is then recycled. Ion exchange resins and solvent extraction systems which may be incorporated into the above process will also be discussed in detail.

Jan 1, 1998

By W. S. Fyfe, G. V. Albino, Ekawat Sitthithaworn

The mineralisation is associated with a composite diorite to quartz monzonite stock of Permian to Triassic age that intrudes Devonian carbonate strata. In the intrusive rocks it is associated with potassic alteration and in the carbonate wall rocks it is hosted by prograde garnet-rich and retrograde epidote-rich skarn. Both porphyry and skarn mineralisation include disseminated and fracture-controlled mineralisation, the latter probably accounting for the major part of the contained minerals. Copper, gold, silver and zinc are strongly positively correlated within the mineralised zones, although there are small differences in metal ratios between the different ore types; molybdenum appears to be negatively correlated with copper and gold. Ore minerals include chalcopyrite, magnetite and pyrite in porphyry-type mineralisation and bornite, chalcopyrite, pyrite, magnetite and lesser haematite in skarn.

Jan 12, 1993

By P Thompson, D Rose, S Yu, W Baum, R Zahn, D Meadows, G Hall

Opportunities to improve mill operations are often cited by mining companies and recognised in many concentrators spread throughout the world but are not always followed through. Copper-molybdenum concentrator operations are challenged by a continuously increasing list of detrimental factors. They include deteriorating ore characteristics and poor mineralogical information on deeper ores, water quality and availability, steel wear and increasing power costs. Rapidly declining productivity, sustainability and environmental factors, new taxes, mining industry global skill set shortages, to name but a few, must also be considered. The days of maximising value by just increasing production or simple cost cutting are fading and will not be sufficient for the future of more complex mineral processing operations and larger plants. As a result, FLSmidth’s Global Process Technology Group has launched a plant and process optimisation program based on mineralogical and metallurgical surveys.Copper-molybdenum concentrators of the future need a paradigm shift. Modelling and simulation tools alone are not the answer to complex mineralogical associations and larger scale equipment and operations. Detailed area by area review aimed at identifying process constraints in terms of equipment and operating practice in conjunction with robust diagnostic work, ie mineralogical and metallurgical characterisation, are mandatory for process optimisation. The navigational beacon, for all operations who want to remain competitive, will be ‘process efficiency’. Since the use of automated mineral analysers has become more readily available and reliable, mineralogical-metallurgical plant surveys have proven themselves as the most efficient tool for optimising plant performance.The following paper will illustrate how operations can move toward eliminating under-performing flow sheets, fixing run-away flotation conditions, maximising individual process equipment performance, preventing shortcuts in Bond work index testing which may cause the failure of comminution models and minimising costly and time-consuming ‘debottlenecking’. As previous plant surveys have shown, the net project cash flow can be increased by 50 to 200 per cent if a combination of robust ore characterisation and plant process mineralogy is performed. The combined plant survey approach has been demonstrated successfully in a multitude of operations. This includes the use of key ore types in concert with statistically sampled process streams, bench-scale metallurgical tests and automated mineral analysis on a size-by-size basis.CITATION:Meadows, D, Hall, G, Rose, D, Thompson, P, Baum, W, Zahn, R and Yu, S, 2014. Copper-molybdenum concentrator surveys using mineralogy and process benchmarking tools to improve overall plant performance, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 315–322 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

By Leonard Larson

DURING several years immediately preceding the adoption of wet-charge smelting at McGill, various necessary conditions affecting this procedure, such as plant rearrangement and the metallurgical nature of the concentrate produced at the concentrator, were fully considered. Wet-charge smelting is not altogether novel to the recently reconstructed smelter at McGill. A number of years prior to this improvement, when flotation methods were initiated at the concentrator, large tonnages of wet and (at that time) sticky flotation concentrate were charged directly into the reverberatory furnaces, because of the difficulty in handling the entire tonnage in the roasters. The plant had excess reverberatory units and this surplus furnace capacity was utilized to smelt raw flotation concentrate. During the years 1927 and 1928, a number of actual operating furnace test runs were made to determine in principle the feasibility of wet-smelting operation. Results were encouraging, and in November 1932 the roasting plant was shut down. For some time thereafter the wet feed was delivered directly to the reverberatory side-charging calcine hoppers already in place. Feeding the wet material through the old calcine hoppers and charge holes into the furnace involved considerable labor, but under the conditions existing at that time it was desirable to continue wet smelting even under the handicap of this laborious procedure. The smelting plant during this period was on a curtailed production basis, operating only about 15 days per month. The alternate starting up and shutting down of the roasting plant presented a difficult problem and charging the wet material directly to the furnace entirely eliminated this difficulty. The roasting plant was never operated thereafter. EARLY PROCEDURE Fig. 1 shows the general layout of the smelting plant before the smelter was remodeled. Under this arrangement, all concentrates and fluxes except converter flux were delivered in cars to wooden storage bins, which paralleled the roaster building. The material to be roasted was dis-

Jan 1, 1938

By Leonard Larson

DURING several years immediately preceding the adoption of wet-charge smelting at McGill, various necessary conditions affecting this procedure, such as plant rearrangement and the metallurgical nature of the concentrate produced at the concentrator, were fully considered. Wet-charge smelting is not altogether novel to the recently recon-structed smelter at McGill. A number of years prior to this improve-ment, when flotation methods were initiated at the concentrator, large tonnages of wet and (at that time) sticky flotation concentrate were charged directly into the reverberatory furnaces, because of the difficulty in handling the entire tonnage in the roasters. The plant had excess reverberatory units and this surplus furnace capacity was utilized to smelt raw flotation concentrate. During the years 1927 and 1928, a number of actual operating furnace test runs were made to determine in principle the feasibility of wet-smelt-ing operation. Results were encouraging, and in November 1932 the roasting plant was shut clown. For some time thereafter the wet feed was delivered directly to the reverberatory side-charging calcine hoppers already in place. Feeding the wet material through the old calcine hoppers and charge holes into the furnace involved considerable labor, but under the conditions existing at that time it was desirable to continue wet smelting even under the handicap of this laborious procedure. The smelting plant during this period was on a curtailed production basis, operating only about 15 days per month. The alternate starting up and shutting clown of the roasting plant presented a difficult problem and charging the wet material directly to the furnace entirely eliminated this difficulty. The roasting plant was never operated thereafter.

Jan 1, 1938

By R. C. Wallace

Introduction At a time when the mineral-bearing areas in Manitoba are being fi very keenly explored because of the major developments which are now taking place in that Province, it may be timely to give an analysis of the mineralization in the two major types of deposits, with a view particularly to assist in the intensive prospecting to which the mineral-bearing areas in the Province are now to be subjected. As is already well known, there are two principal types of mineral deposits in the territory- copper-zinc and gold; and attention will be confined almost exclusively in this paper to these two types, with some indication of such modifications of the main types as may be found to be of economic value. The details of the paper are drawn from what is yet a small part of the area of the Province available for prospecting. Enough is now known of the types of mineral deposits which are described in the following pages to justify the expectation that these types may serve as criteria in the search in the newer districts. Doubtless new types will also be found as investigations proceed.

Jan 1, 1928

By Yongfeng Jia

Lime neutralization and co-precipitation of As(V) with iron(III) is the most widely used and most economical method for the removal of arsenic from mineral processing solutions and effluents. However, the arsenate fixation mechanism in the coprecipitated solids is not fully understood. It was suggested by various researchers to be a ferrihydrite precipitation process during which arsenate is removed by adsorption on ferrihydrite. In this study, XRD, EXAFS and microprobe were employed for the speciation of arsenic and the effect of various factors on arsenic solubility was also investigated. The results indicate that coprecipitation is a more complicated process than simple adsorption resulting in lower arsenic solubilities. Poorly crystalline ferric arsenate was observed for the first time to be present in the coprecipitated solids. In case of lime as base, there exists some type of Ca-Fe-arsenate association. A general term ?coprecipitate? is suggested to describe the precipitated arsenic-bearing solids instead of those widely used specific terms ?arsenical ferrihydrite? or ?arsenic-bearing ferrihydrite?.

Jan 1, 2005

By G N. Musafer, M H. Thompson, R C. Wolff, E Kozan

The most important aspect of modelling a geological variable, such as metal grade, is the spatial correlation. Spatial correlation describes the relationship between realisations of a geological variable sampled at different locations. Any method for spatially modelling such a variable should be capable of accurately estimating the true spatial correlation. Conventional kriged models are the most commonly used in mining for estimating grade or other variables at unsampled locations, and these models use the variogram or covariance function to model the spatial correlations in the process of estimation. However, this usage assumes the relationships of the observations of the variable of interest at nearby locations are only influenced by the vector distance between the locations. This means that these models assume linear spatial correlation of grade. In reality, the relationship with an observation of grade at a nearby location may be influenced by both distance between the locations and the value of the observations (ie non-linear spatial correlation, such as may exist for variables of interest in geometallurgy). Hence this may lead to inaccurate estimation of the ore reserve if a kriged model is used for estimating grade of unsampled locations when non-linear spatial correlation is present. Copula-based methods, which are widely used in financial and actuarial modelling to quantify the non-linear dependence structures, may offer a solution. This method was introduced by Bßrdossy and Li (2008) to geostatistical modelling to quantify the non-linear spatial dependence structure in a groundwater quality measurement network. Their copula-based spatial modelling is applied in this research paper to estimate the grade of 3D blocks. Furthermore, real-world mining data is used to validate this model. These copula-based grade estimates are compared with the results of conventional ordinary and lognormal kriging to present the reliability of this method.CITATION:Musafer, G N, Thompson, M H, Kozan, E and Wolff, R C, 2013. Copula-based spatial modelling of geometallurgical variables, in Proceedings The Second AusIMM International Geometallurgy Conference (GeoMet) 2013 , pp 239-246 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 29, 2013

By A. Sutherland Brown

"The Red Rose Mine is in the Rocher Deboule Range, 8 miles south of Hazelton. The mill camp (elevation 4,000 feet) is on Red Rose Creek, 11 miles by road from Skeena Crossing. The mine camp is one mile northeast and 1,600 feet above the mill camp on a spur of Red Rose Peak.The Red Rose mine was described in the Jubilee Volume, Structural Geology of Canadian Ore Deposits (Davis, pp. 129-131), but regional geology and mine development have progressed to a stage at which a more complete description is possible.The property was first located in 1912, and for the next 7 years was prospected actively for gold and copper on the southern extension of the Red Rose shear. Scheelite was first discovered in the vein in the northern part of the shear in 1923. The Consolidated Mining and Smelting Company of Canada, Limited optioned the property in 1939, and after a drilling program in 1940 the present mine development was initiated in the vein in the northern part of the shear. A 75-ton-per-day mill started operating early in 1942 and continued until 1943. The property was inactive until 195 l, when Western Tungsten Copper Mines Limited leased the property from The Consolidated Mining and Smelting Company of Canada, Limited. A mill, which later was enlarged to 140 tons per day, started operations at the end of 1951. The mine and mill were closed in December, 1954.The mine has twelve levels and sub1evels, and most of the workings are in the plane of the vein (see Figure 2a). The level interval is not constant, but between the lower four levels is 87 feet (100 feet in the plane of the vein). Access is provided by four adits; the 800 (elevation 5,659 feet), 600 (5,920 feet), 300 (6, 133 feet) and 200 levels (6,237 feet). The 600 level is the main haulage, and from it the ore is transported to the mill by aerial tram. From this level an inclined shaft extends in the vein to the lowest level (I, 100)."

Jan 1, 1949