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By A. C. Hawken, L. M. Strah

"The Report outlines the Conversion to Rubber Mill Liners and steel grinding balls from the old style Cobalt pocket liners and flint pebbles which had been used in 5'x22' Fraser-Chalmers tube mills for over fifty years. An outline of the past history and some of the plant modifications are mentioned as well.The details pertaining to considerations, modifications and repairs carried out on the mills are described. Factors influencing the rubber liner design are outlined along with the description of the rubber liner design.The performance and advantages to date are given and concludes with brief summary.HISTORYThe original amalgamation-cyanidation mill of Dome Mines Limited went into operation in 1912. Milling has been continuous since 1912, except for two periods; one of 17 months commencing in December 1917, and the other of about 12 months, when fire destroyed the plant in November 1929. Initially the capacity of the mill was 350 tons a day. In 1916 it was raised to 750 tons a day, following which there was a gradual expansion to 1,500 tons a day."

Jan 1, 1977

By Maurice Lauzon

"Aerial surveys of the open pits have been conducted regularly at Lake Asbestos of Quebec, Ltd. -a subsidiary of American Smelting and Re-fining Company -since 1958. Ground controls are supplied by Lake Asbestos, and a topographic map of the pits at a scale of 1 in. to 50 ft., with 2-foot contour intervals, is obtained by aerial photography. The information from the topographic map is then transferred by the Lake Asbestos engineering staff to sections from which the volumes removed are calculated. Considering the size and complexity of the operation, this system has proved to be the cheapest, fastest and most accurate way of determining the quantities of rock and overburden removed. More than 11,000,000 tons of ore and waste are mined each year in three different pits. Further-more, the nature, shape and extent of the orebodies make it necessary to mine on many (up to eight) benches of a pit during a single year. In addition, there are sections of the pits where it is impossible or un-safe to survey the crest of some benches by standard ground methods. For these reasons, the tonnages obtained by truck count are checked and adjusted by aerial survey. These surveys are now taking place every six months. With this system, it is possible for a two-man surveying crew to look after all field work in the mine area."

Jan 1, 1967

By L G. White

During 1958, prospecting parties working for the Mackenzie Syndicate in the Flat River area of the Mackenzie mountains, Northwest Territories, identified tungsten mineralization and located the property now being developed by the Canada Tungsten Mining Corporation Ltd. Scheelite mineralization, associated with pyrrhotite and minor chalcopyrite, occurs in lime silicate skarn replacement zones within a limestone-chert-argillite series intruded by granitic dykes originating from the ? main batholithic mass underlying the sedimentary beds. The results of surface outcrop sampling and 15,000 feet of grid-pattern diamond drilling have been combined to outline an orebody estimated to contain 1.32 million tons, grading 2.51 per cent WO after applying a 10 per cent dilution factor for mining purposes. T HE property of Canada Tungsten Mining Corporation Ltd. is located in the upper Flat River area of the Mackenzie mountains, Northwest Territories. Skarn-type replacement deposits containing scheelite mineralization are .situated between elevations of 4,500 and 5,500 feet above sea level along the east flank of the mountain range forming the boundary between the Yukon Territory and the Northwest Territories. The nearest supply center for the area is Watson Lake, Yukon Territory, located at Milepost 635 on the Alaska Highway.

Jan 1, 1963

By W. H. MacLean, P. J. MacGeehan, A. J. Bonenfant

"The Norita deposit is an Archean volcanogenic-exhalative Fe-Zn- Cu sulphide accumulation of the distal type, consisting of six overlapping stratiform massive sulphide lenses, interbedded with exhalative cherts and thin pillow basalt flows, which formed down-topographic-slope from a mineralized hydrothermal discharge zone in the Norita rhyolite. The sulphide lenses are of two main types: channel-fill deposits such as the Main Zone and Main Zone West, composed of intricately layered pyrite-sphalerite-chert, which formed by the dynamic down-slope transport and re-deposition of sulphide exhaled from the vent zone; and stratiform lenses such as the Cu-rich A-Zone and A-Zone West, which formed even farther down-slope, in basinal depressions flanking the dome. Although rotated into a vertical position by folding, the sulphide lenses are otherwise essentially undeformed and contain a wide range of preserved primary sedimentary structures.Of prime importance in exploration has been the recognition that the formation of the distal sulphide lenses was controlled by submarine paleotopography, and that it is possible to approximate the sea-floor paleotopography using variations in the thickness and strike extent of the interbedded volcanic rocks. This has proved to be a major and successful exploration tool in the search for further sulphide mineralization in the mine, and offers prospects for use elsewhere in the district."

Jan 1, 1981

By G. Barr

"The Polymetallic Twin Metals Project deposits in Northern Minnesota are being developed by Twin Metals Minnesota. Amongst the flowsheet alternatives being examined at a prefeasibility study level is one involving flotation of a bulk concentrate followed by high temperature autoclaving of the concentrate using the PlatsolTM Process to solubilize all the valuable metals in the concentrate (Cu, Ni, Pt, Pd and Au). Autoclave operating conditions have been established in which recoveries of all the pay metals are in the 92 to 99% range, and the autoclave residence time has been reduced to an hour. The stability of the precious metals in solution after autoclaving has been assessed, as well as the impact of internally recycling process liquors to the feed to the autoclave.INTRODUCTIONPlatinum group metals (PGMs; platinum, palladium, rhodium, ruthenium, iridium and osmium) are traditionally recovered from their ores by production of a high-grade concentrate by flotation and/or gravity concentration followed by smelting of the concentrate. Concentrates are smelted to remove nonsulphide minerals and converted to remove iron and sulphur. The resulting high PGM grade matte is processed hydrometallurgically to leach the base metals, leaving a residue with a very high grade (20-40%) PGM concentrate for final refining to individual PGMs. Due to the complexity of this process and its high capital cost, this process can only be applied economically to high grade PGM ores. Lower grade PGM concentrates might be accepted as copper and nickel concentrates but may not yield attractive payment terms for the PGMs. The mineralized Duluth Gabbro region of Minnesota hosts several low grade base and PGM deposits and Twin Metals Minnesota is working to develop the Twin Metals Project deposits. Earlier studies at both pre-feasibility and feasibility level, by other mining companies, have failed to yield an economically viable project based on the traditional process. It is possible to produce a saleable copper concentrate from these ores, but economic recovery of nickel and PGMs into saleable concentrates proved elusive.In the late 1990’s and early 2000’s the Platsol™ process was developed at SGS in Lakefield, Ontario (at the time Lakefield Research Limited) for the NorthMet deposit (also within the Duluth Gabbro) owned by PolyMet Mining. The process developed was a high temperature chloride-assisted pressure oxidation process capable of dissolving simultaneously the base metals (Cu, Ni, Co, Zn), gold and the PGMs."

Jan 1, 2012

By J. M. Goris

"Researchers at the U.S. Bureau of Mines are conducting research on cable-type ground supports to assess their material and support properties and to provide design criteria for implementing cable supports as a workable roof control system under various underground mining conditions. Much of this work involves laboratory assessment of the material and support properties of cable supports using both conventional and birdcage steel cables. Included in this assessment are single and double cables, effects of changes in embedment length on support strengths, water-cement ratios, use of breather tubes, and pumpability and water-bleeding properties of grouts.IntroductionSince cable supports were introduced in the early 1970s, considerable research has been conducted in laboratories and at mine sites to understand their behaviour. The U.S. Bureau of Mines began a laboratory evaluation project in 1985 at the Spokane Research Center (SRC) to study the behaviour of cables supports under uniaxial loads so that the effects of various cable support components, such as grout and breather tubes, could be identified. Other objectives were to compare the behaviour of various types of cables and grouts and to collect meaningful data for numerical analyses. Pull tests were conducted to determine how cables supports reacted under uniaxial loads and how they failed when ultimate loads were exceeded. Such data are essential for the development of numerical models for analyzing rock mass behaviour when the rock is supported with cables.Test ProgramBecause each cable support can be composed of either single or double cables as well as a breather tube, pull tests were designed to include a variety of samples containing these components. A series of samples containing a single conventional cable, a neat cement grout having a water-cement ratio of 0.45, and no chemical additives was selected as a standard (Table I). Then one component in the system, such as the presence of a breather tube, was changed, and results of tests compared to results obtained for the standard samples to determine what influence that component might have on the behaviour of the support system. If significant changes in test data occurred, they could be attributed to that one component. To control the quality of the samples, the physical condition of the cables and tubes were monitored closely, and compressive and tensile strengths as well as flow properties were determined for the grout used in each test series."

Jan 1, 1991

By M. Askari, A. Ahmadi

"This study was conducted to establish the possible application of bioleaching in the pilot scale to recover valuable metals from the sulfidic tailings of Golgohar Iron Mine (Kerman, Iran). Using bioleaching processes, solvent extraction and sulfide precipitation recovery of valuable metals including, copper, cobalt and nickel from the tailing was investigated. The results showed that the optimum pH and temperature conditions relative to the bioleaching step was 1.5 and 45°C. The extraction of the metals from the low grade polymetal resource with complex mineralogy is very important from economic, environmental and technical point of views. Moreover, the proposed method has some advantages from energy saving, tailing management and creating added value from waste materials aspects.INTRODUCTION The growing demand for base metals and strategic such as: copper, nickel and cobalt as well as high-grade resource depletion has caused them, the mining industry to extract metals from low-grade resources and driven by complex mineralogy (Xie et al., 2005). Economic extraction of metals with low grade sources requires low-cost and efficient processing methods. A source with enormous potential for metals in future, secondary sources, such as: tailings mineral industries which the lack of appropriate technology and the high cost of processing weren't recoverable with previous procedures in the past. But it is possible that with develop new and efficient technologies have the economic exploitation ability. Recovery of metals from tailings resources is very important not only in terms of precious metals but also in terms of reducing environmental problems (due to heavy metals). The metals of copper, nickel and cobalt are mainly like sulfide compounds in the nature, usually for their recovery, first they make ore with flotation method to obtain the sulfide concentrate that containing valuable metals. then for the metal production, it will be melt and purification. Due to the environmental aspects, technical problems in the recovery of base metals from low grade ore and mixes, the presence of small reserves, in the past three decades they pay much attention to the use of hydrometallurgical processes for the recovery of base metals from these resources. These processes are classified according to the solver agent to leaching sulfates, chlorides, nitrates and amines (Hackl, 1995). Early, chloride leaching under oxidation condition in acidic or alkaline solutions was recognized the best process to produce sulfide concentrates. But then the basic processes sulfates was considered as the superior process because of being easier, a better understanding of the chemistry of leaching and recovery of base metals, especially copper from the environment easier sulfate solution with an organic solvent extraction and electro winning. The bioleaching is shown its ability on a commercial scale of low-grade sulfide minerals and ore that because of environmental reasons, the possibility of melting is not possible to them. The existing industrial base factories are based on primarily on the bioleaching and pressure leaching and recovery of cathode copper sulfate solution by extraction with organic solvents (SX) and electrowinning (EW) (Wang, 2005). Preliminary analysis has shown that valuable elements such as copper, nickel and cobalt are in the float flotation unit. Hence the plan to extract these metals from the wet tailing stream has been done."

Jan 1, 2017

By G. H. Robbins

"Originally presented at TORONTO 94, CIM-AGM, this paper has been updated to reflect the present status of business development for theSO2/AIR process. The paper reviews the back-ground leading to the invention, discusses the process chemistry and historically tracks the commercial development. Engineering and cost considerations are presented.IntroductionThe year 1994 was particularly significant in terms of the technical and business development of the INCO SO2/AIR cyanide destruction process. The year marked a decade of service to the world mining community since INCO acquired the Canadian patent in 1984. In that time period, the technology had become widely recognized as the state-of-the-art process in the field of cyanide detoxification and today we enjoy the distinction of having licensed the process at over 50 project sites worldwide. The application for the process has mostly been in the mining industry in precious metal recovery plants where it offers treatment for a wide range of waste streams such as: CIP/CIL slurries, Merrill-Crowe barren bleed solutions and/or re-pulped filter cakes, tailings pond water recycle or discharge, make-up water to flotation circuits and feed to underground backfill plants. The process also applies to the treatment of high cyanide rinsewaters from metal plating shops. More recently the process is being applied to the rinsing and abandonment of heap leach pads where it has proven to be a very cost-effective application of the INCO process. The distinct advantages that have led to the success of the process are the ability to work on slurries as well as solutions and the significantly low cost of sulphur dioxide products. The process is safe and easy to manage and uses all forms of SO2(liquid, gas or salts) which gives the process tremendous versatility with respect to geographic application. Having more than ten years of service"

Jan 1, 1996

By D. Van Tonder, C. Edwards

ABSTRACT Solvent extraction is used in many acid leach uranium circuits to upgrade and purify the uranium-bearing solution prior to final product precipitation. A number of strip options are available to recover the uranium from the loaded organic. This paper compares seven of these options on the basis of typical capital and operating costs for the solvent extraction circuit and downstream uranium processing and also considers other related issues. The aim is to provide an objective comparative assessment of these options that could be used as a qualitative guide in process selection or improvement.

Jan 1, 2013

By L. R. Hougen

The Falconbridge Matte Leach Process yields high-purity nickel chloride by selectively dissolving nickel from copper-nickel converter matte with hydrochloric acid , purifying the solution by solvent extraction and crystallizing NiCl2 4H2O by further acidifying the purified leach liquor. In the Kristiansand Matte Leach Plant, a granular oxide is produced by high-temperature hydrolysis of the nickel chloride for subsequent reduction to a high-purity granular metal product, designated NICKEL 98. This paper describes the investigation of on alternate route to processing nickel chloride in which high-quality cathode nickel is produced by electrowinning from all- chloride electrolyte. The investigation was made in a pilot circuit containing two production-size electrowinning tanks operated in conjunction with the commercial-scale leach plant.

Jan 1, 1977

By T. R. C. Aston, C. M. K. Yuen, J. L. Gilby

"The access tunnels for the Donkin-Morien Mine in Cape Breton, Nova Scotia, comprise two parallel tunnels driven through a sedimentary sequence of sandstone, siltstone and mudstone to intersect the Harbour Coal Seam under the Atlantic Ocean, approximately 3.5 km offshore (Fig. 1). The No. 2 tunnel was driven a total distance of 3579 m using a 7.6 m diameter Loyat full-face tunnel boring machine (TBM) and was supported by steel sets. The No. 3 tunnel was initially driven a distance of 1027 m using drill and blast methods (dimensions 7.6 m wide by 5.3 m high) supported by rock bolts and shotcrete. This tunnel was then completed (front March 1986 to February 1987) using the tunnel boring machine to a distance of 3573 m.During the period May 1983 to December 1984, a major geotechnical research program was undertaken in the Donkin-Morien tunnels by Golder Associates in conjunction w,ith CANMET and the Cape Breton Development Corporation (Yuen et al., 1985). The research program included the monitoring of the performance of the TBM and drill and blast drivages, and the study of the thrusting characteristics and rock-interaction of the TBM excavation. The main emphasis of the research program was on the collection of field data which was to be made available in the form of a comprehensive report for the future use of others who might wish to carry out a detailed interpretation and analysis. The information gained from the program was also used during the tunnel"

Jan 1, 1988

By R. L. Bailey

"The Acheson D-3 Pool was discovered in August 1950, and contains 88 producing oil wells on 40 acre spacing. Initial oil-in-place was 118 million stock tank barrels and initial well potentials averaged 2,000 barrels per day.In early 1956 studies of hydraulic interference between the Acheson and Leduc D-3 Pools were conducted on the Carter Petroleum Reservoir Analyzer. These studies indicated that the natural water drive to Acheson would be incapable of replacing reasonable withdrawal rates over the life of the pool. To avoid shrinkage losses a scheme of pressure maintenance by water• injection was therefore instituted in March 1957.This pressure maintenance scheme is notable because pressure is being effectively maintained by high volume injection through one well located on the flank of the pool. Close control on reservoir behaviour is maintained by regular bottom-hole pres-sure surveys and observation of water cuts on producing wells. No evidence of pressure gradients or channelling of the injected water has been observed to date, and the water-oil contact has remained level over the whole field. It is possible to control the average reservoir pressure within very close limits by varying the rate of fresh water injection.Ultimate recovery estimated from the rise of the oil-water contact and calculated water influx to December 31, 1958, is 335 barrels per acre foot, or a total of 91.4 million barrels from the pool. This compares to 217 barrels Per acre foot or a total of 59.4 million barrels indicated from laboratory test data."

Jan 1, 1961

By E. Poiate, P. R. Cabral, A. V. N. Santos, P. V. M. Costa, C. S. Amaral, A. M. Costa

This paper discusses the design of underground salt caverns, opened by solution mining, for the production of brine and storage of natural gas under high pressures. The caverns are located in the salt basin in Maceio City, Alagoas State, at the northeast region of Brazil. Special attention is given to the geomechanical structural design of the caverns using computer codes specially developed for the simulation of excavations in salt rock formations, considering validated constitutive creep laws under different initial stress states and temperatures. The behavior of sedimentary rocks intercalations presented in the salt layer is simulated using elastic-plastic models. The excavation of the caverns by solution mining is also simulated starting from a geostatic initial equilibrium state of stresses, with a unique formulation only implemented in these computer codes, without the need of the application of boundary stresses to the model in order to develop the initial state of stresses. These computer codes have being used in the design of several clusters of caverns for brine production used as raw material of Poly Vinyl Chloride (PVC) of the Chemical Industryin Brazil. These computer codes and methodology were also applied to the design of the underground conventional potash mine in the state of Sergipe, overlying tachyhydrite and carnallite with very high creep strain rate. They have also been used in the design of the pre-salt oil wells through thick layers of stratified salt rock in Santos Basin, for the exploration and production of the pre-salt reservoirs in Brazil. The constitutive creep equations implemented in the codes are based on a double mechanism creep law. The creep constants adopted in the simulation are taken from a previous investigation carried out on an experimental panel of the potash mine, considering the stratified geology with different salt rocks of the basin where the mine is located. These constants were also obtained by lab tests under different confined and differential stresses and temperatures.

Jan 1, 2015

By C. J. Newman, K. Molnar, A. G. Storey, G. Macfarlane

"The Kidd Creek copper smelter started operation in June 1981 using the Mitsubishi Continuous Smelting Process. A steady increase in plant production rates has been demonstrated, and the annual tonnage of concentrate smelted is now double the original design rate. This paper summarizes recent modifications made to increase concentrate feed rates, increase on-line time, reduce copper losses and control the impurity level in anode copper. It includes improvements made to furnace design, improvements to reduce downtime f or boiler cleaning, and improvements made to the acid plant to allow handling of increased gas flowrates and higher gas strengths while maintaining low acid plant emission rates. IntroductionThe Kidd Creek Division of Fal conbridge Limited operates a base-metal mining and metallurgical complex, located in Timmins, Ontario, which includes an underground mine, a 13 500 tonne/day concentrator, a 130000 tonne/year zinc plant and a copper smelter and refinery, The copper smelter and refinery were commissioned in 1981, with an installed capacity of 59 000 tonnes/ year of cathode. The smelter, utilizing the continuous smelting technology of the Mitsubishi Process, has subsequently been expanded in production capacity, and in 1991 produced 113 500 tonnes of blister copper The capacity of the refinery, which uses Onahama No. 3 tankhouse technology, has been increased to 96 000 tonnes/year of cathode. The capacity increase at the smelter has been realized through the input of additional oxygen to both the smelting and converting furnaces, without increasing the process off-gas flowrate through the associated waste heat boilers.In this paper, a brief review of smelter operation and modifications in the period 1982 to 1987 is presented, together with a more detailed discussion of operation during 1988 to 1991, and a discussion of future planned improvements."

Jan 1, 1992

By Fan Delian, Ye Jie, Lui Tiebing

The lower and the upper parts of the Doushantuo Fonnation, Hubei Province, China, contain silicate-type K-rich beds (illite and feldspar), and intraclastic phosphorites, as well as Ag-V and V deposits in a "Black Shale Series". The K-rich beds fonned in an anoxic basin whereas the phosphorites fonned in a more oxic and higher energy depositional environment. The Black Shale Series occurs in the lower and upper parts of the Doushantuo Fonnation; that in the upper part consists of black illite shale and dark gray dolostone, and contains Se-rich Ag and V deposits. These deposits are of more industrial significance than the previously discovered Ni-Mo polyelement sulfide deposits which occur in the Lower Cambrian Black Shales Series of South China. Studies of lithotype, sedimentary sequence, and geochemistry of the Doushantuo Fonnation have led to more attention being focused on the uppennost Black Shale Series and its Ag-V and V deposits. There are three ore types: black, Ag-V-rich, illite shale; black, V-rich, illite shale; and dark gray, Ag-Vrich, fmely crystalline dolostone. The first two ore types contain 3070 to 5% Corg and the last, 1.25%. The Ag-bearing minerals are naumannite, argyrodite, aguilarite, acanthite, and native silver that occur as inclusions in pyrite associated with barite. Vanadium substitutes for AI in illite and is also related to organic-matter contents. Because (1) the pre-Sinian Kongling Group contains one to three times greater than the Clark Ag value; (2) black, illite shale has a negative Ce anomaly; and (3) there are large occurrences of black, K-rich, tuffaceous, fme siltstone, the sources of Ag are thought to include subaerial and submarine volcanic materials. It is proposed that Ag, V, and Se were absorbed by clay particles and organic matter during syngenetic deposition in an anoxic environment; Ag subsequently was reconcentrated into sulfides and selenides, whereas V was incorporated structurally into illite during early and late diagenesis. These types of Black Shale Series-hosted Ag-V and V deposits are, therefore, considered to be sedimentary-diagenetic in origin.

Jan 1, 1992

By F. Vasilikou

Commercial deployment of Carbon Capture Utilization and Storage (CCUS) requires field testing of a scale that can stress the geologic reservoirs. One such reservoir of interest consists of unminable coal seams that exhibit favorable characteristics and depositional environments and lower pressure and temperature than other, deeper, reservoirs. Such conditions can reduce compression costs while utilizing the action of adsorption that offers a more effective carbon dioxide (CO2) bonding than free storage or solution. To ensure the success of such tests, a number of parameters should be accurately determined, such as the seam geometry and stratigraphy, coal porosity and permeability parameters as well as optimum injection conditions. In essence the CO2 injection model should be calibrated for the reservoir characteristics using all available data. History matching calibrations are based on gas and water production from existing wells prior to injection. This paper will present reservoir models that were developed for a pilot--scale 907 tonnes (1000 tons) CO2 injection test that was performed in 2009 through one legacy coalbed methane production well in Russell County, Virginia, USA. The model incorporates a number of individual coal seams about 0.3 m (1ft) in thickness located at depths ranging from 300 to 700 m (1,000 to 2,200 feet). Model calibration was performed through history matching to prior production data, and subsequently the model was utilized to develop different injection scenarios. The developed model was used after injection to calculate CO2 plume distribution patterns that were monitored at the injection test. The paper will present model data and assumptions with a special emphasis on the effect of hydraulic fractures and the skin factor to the coalbed methane (CBM) production model.

Aug 1, 2013

By Donald Brisebois

The mining of high-grade, narrow vein deposits is an important field of activity in the precious metal mining sector. The principle factor that has undermined the profitability and effectiveness of mining such ore zones is the substantial dilution that occurs when blasting with explosives during extraction. In order to minimise dilution, the Thermal Fragmentation Mining Method enables the operator to extract a narrow mineralised corridor, 50 cm to 1 metre wide (according to the width of the vein), between two sub-level drifts. By inserting a strong burner powered by diesel fuel and compressed air into a pilot hole previously drilled directly into the vein, a thermal reaction is created, spalling the rock and enlarging the hole to 80 cm in diameter. The remaining ore between the thermal holes is broken loose using low powered explosives, leaving the waste walls intact. This patented method produces highly concentrated ore, resulting in 400% - 500% less dilution when compared to conventional mining methods. The mining method reduces the environmental affects of mining operations since much smaller quantities of rock are displaced, stockpiled, and treated using chemical agents. The fully mechanised equipment operated by a 2-person team (1 thermal fragmentation operator, 1 drilling operator) maximises the effectiveness of skilled personnel, and increases productivity and safety. The Thermal Fragmentation is currently employed in 3 mining operations in North America.

May 1, 2010

By C. H. B. Leitch

"The Bralorne-Pioneer deposit produced more gold than any other deposit in British Columbia over its 70 years of operation (130 kg or 4 million oz of Au, from 7 million tonnes of 18 glt ore). Mineralization is related to a suite of late Cretaceous albitite dykes of 85 to 90 Ma age. It thus occurred long after and is genetically unrelated to the Early Permian (275 Ma) Bralorne diorite and soda granite.The major veins strike about 110 degrees and dip north at 70 degrees, with slickensides plunging 45 degrees east that indicate last movement was reverse. Major ore shoots in the veins occupy somewhat less than 20% of the vein and plunge steeply west, roughly perpendicular to the slickensides. The structures that contain the veins probably formed at around 95 to 90 Ma during a strike-slip episode with easterly directed maximum compression.Disseminated pyrite, pyrrhotite, and minor chalcopyrite occur within envelopes up to several metres from the veins. Arsenopyrite is confined to vein selvages. Local sphalerite and galena appear to correlate with gold-rich portions of the veins. Traces of tetrahedrite and stibnite have been observed but tellurides have not. Gold is found principally as thin smeared flakes of the native metal in the black sulphidic septae of the strongly ribboned veins. Hydrothermal alteration envelopes around the veins grade outwards from intensely foliated quartz - ankerite - sercite (±fuchsite) to less sheared calcite - chlorite - albite to unsheared epidote -calcite.Open-space vein filling textures found in the deep (mesothermal) environment at Bralorne can be explained by a fault- valve model. High-angle reverse faults may have acted as fluid-pressure activated valves on geopressured reservoirs, causing large cyclic fluctuations in fluid pressure that promoted mineral deposition in an episodic fashion . The faults that were mineralized were reactivated sub-vertical former strike slip faults, oriented at a high angle to the northeast-directed stress in a compressive regime."

Jan 1, 1990

By W. N. Wallinger

The mine and concentrator of Cyprus Anvil Mining Corporation are situated in south-central Yukon, some 120 miles north of Whitehorse. The large lead-zinc concentrator employs conventional crushing and grinding , followed by differential flotation, to produce selective lead and zinc concentrates, as well as a mixed lead-zinc (bulk) concentrate. Since the commencement of operations in 1969, the concentrator has undergone expansions and modifications which have increased the capacity to 10,000 tons per day, and afforded some metallurgical improvements.

Jan 1, 1978

By E. W. Shaw

The past year (1950) has been a big one for the oil industry in Canada, with new records being set in almost every phase of exploration and development activity. New highs were reached in acreage holdings, in geophysical activity, in wildcat drilling, and in field development and production. The only item that did not show a substantial increase over recent years was the amount of reserves added by new discoveries. This was not due to a decrease in the number of new discoveries, however, hut rather to the fact that most of the discoveries were made during the later months in the year and could not be fully evaluated by the year end. The foregoing remarks pertain mainly to activity in Western Canada. In Eastern Canada, operations continued on a modest scale. with exploration drilling down slightly from the previous year but with development drilling and production at about the same level. Returning to Western Canada, it is observed that the general exploration play was one of expansion and intensification. The issuing of new petroleum and natural gas regulations by the province of British Columbia precipitated a moderate entry of companies into the northeastern part of the Province.

Jan 1, 1951