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By Ian Hattie, Mark R. Wilson, T. Kurtis Kyser, Brian W. Janser

Petrographic, chemical and isotopic compo itions of auriferous and non-auriferous quartz vei ns from the Sta r Lake and Flin Flin-Amisk Lake areas of northern Saskatchewan indicate diver e origins fo r the fluids associated with gold mineralization. In the Sta r Lake area, the metavolcanic , Star Lake pl uton, and heared mafic dikes have retained their original igneous 81 80 values. The 81 80 value of quartz from mi neralized veins vary by I per mi! and range from+I0 to+IJ. The only quartz ample having a 81 80 value outside th is range came from a vein not associated wi th a dike or gold mineralization . Quartz from the Star Lake pluton has identical 8180 value to the qua rtz from mi neralized vei ns which suggests that the vei ns and the pluton are related to imilar proces es. The diffe rence in 81 80 values between mi neralized and barren vein may erve a a potential exploration tool in this area. Because o34S val ues of a limi ted number of amples of pyrite from gold-bearing quartz vei ns at Star Lake correlate directly with gold content. the gold may have been syngenetic with the su lphur and the 834 value may be useful for exploration . 8180 values of co-existing quartz, calcite, and amphibole in the veins from Star Lake indicate that isotopic disequilibrium among the minerals is common, but the val ues also record a high temperature of formation (ea . 500°C) involvi ng a component of magmatic fl uid. TheoD value of hydrou mineral from Star Lake retain their original igneous values but whole rock values are generally much lower, an effect attributed to interaction of the minerals with a late meteoric water. In the Fl in Flin-Amisk Lake area, auriferou and barren quartz veins formed from a mixture of hydrothermal and metamorph ic flu id at lower temperatures (ea . 350°C) than those at Sta r Lake. Becau e there i no apparent correlation between 8180 values of the quartz with gold content or relative age of the vein , nor between the 834S val ues of pyrite and gold content, table isotopes appear limited as an exploration tool in the Fl in FlonAmisk Lake areas.

Jan 1, 1986

By F J. Kelly

A chemical process, based on laboratory investigations, is proposed for the production of high-purity niobium oxide from a pyrochlore-perovsk1te flotation concentrate produced from ore obtained at the Oka, Quebec property of Columbium Mining Products Limited. The method developed included the decomposition and dissolution of the flotation concentrate with hot, concentrated sulphuric acid. The dissolved niobium was purified and concentrated by a solvent extraction step, using methyl isobutyl ketone as the extractant. The niobium was recovered from the extractant with ammonium fluoride solution from which niobium oxide was precipitated with ammonia. Finally, the filtered and washed precipitate was calcined. Ninety-six per cent of the niobium was recovered in a product analysing 99.9 per cent Nzo.. The reagent cost was in the range of $1.00 to $1.25 per pound of niobium recovered. Introduction OVER the past few years, Columbium Mining Products Limited has been engaged in the development of a niobium-bearing mineral deposit in the vicinity of Oka, Quebec. One phase of this program resulted in a flotation process capable of producing concentrates containing up to 50 per cent Nb205? Flotation pilot-plant

Jan 1, 1965

By B. Jin

Removal of fluoride and chloride by alkaline washing from direct lead smelting dust which originated from an Ausmelt furnace is reported. The results show that the fluoride removal efficiency increases with an increase of alkaline concentration, alkaline washing temperature and time. The L/S ratio has little effect on fluoride and chloride removal efficiency. A new flow sheet for removal of fluoride and chloride from the dust was successfully proposed according to the results. Fluoride and chloride removal efficiencies are more than 53.68% and 63.18% respectively using the flow sheet under the optimum parameters of temperature 338 K, L/S ratio 3:1 and alkaline washing time of 1.5 h. The contents of fluoride and chloride in the alkaline washing residues can meet the standard of zinc purity as hydrometallurgical materials for zinc processing.

Jan 1, 2014

By Dao Dac Tao

The Vietnam coal industry in recent years has achieved the rapid growth. However, in the process of industrialization and modernization of the country, the demand for coal for energy is constantly increasing. This paper presents the Vietnam coal industry in the last stage: the opportunities and challenges facing the coal industry as it strives to become modern, highly effective and capable of meeting the demands of national energy security, all in harmony with socio-economic and environmental needs.

Aug 1, 2013

By Adrian Dance

The Antamina orebody is a geologically complex, polymetallic skarn located in the Peruvian Andes. A range of copper and copper-zinc ores are campaigned separately through the concentrator producing copper and zinc concentrates as well as lead and molybdenum by-products. Ore type classification at Antamina is based on chemical assays; geological or lithological variations are not considered. Consequently, each ore type demonstrates a wide range of concentrator performance. Antamina is currently implementing a system to track ore movements from the shovel face to the concentrator. Relatively sophisticated models of the ROM and coarse ore stockpiles ahead of the mill will be used to account for the variable delay between mining and milling. One of the numerous benefits of ore tracking is the ability to directly associate the measured concentrator performance with a physical location in the orebody. Using existing mine planning software, three-dimensional models can be created of such performance measures as metal recovery, concentrate quality, operating cost and reagent consumption. Such a database of information allows the true value of different ore sources to be determined. The value of an ore source is conventionally measured as metal content (individual or combined ?equivalent?) with a cut-off grade applied to indicate some base level operating cost. However, the true value of an ore source should include dynamic factors like haulage cost, reagent use, smelter penalties and even prevailing metal prices. If an ore source is considered from all perspectives rather than simply its potential to deliver value, the concept of ?ore? becomes more flexible. Any ore source that reveals itself to be unprofitable can be eliminated from the mine plan or possibly blended with others to improve its value. The ability to predict mill performance ? and thereby profitability ? is a powerful tool. In an operation such as Antamina, where managing ore sources, campaigns and the delivery of a variety of concentrates in a narrow timeline to customers is a daily challenge, prediction is the key to success and improved efficiency.

May 1, 2003

By M. O. Makwarela, M. W. Erwee, K. Bisaka

"Energy efficiency in mining and minerals processing forms part of Mintek’s long-term research strategy which includes a specific programme designed at overcoming the challenges faced by the South African pyrometallurgy industry.The electric smelting industry is currently severely constrained by the global economic downturn as well as by the availability and increasing cost of electricity. Efficient utilisation of electricity through energy recovery systems and improved smelting process flowsheets can significantly contribute towards creating sustainable operations.Apart from efficient engineering design and solid process control, smelting of either cold or hot pre-reduced ores has been shown to decrease furnace electricity consumption by more than 40%. Various methods of solid-state reduction tests have been carried out at Mintek in order to select the most appropriate process option in the context of the South African industry. This paper summarises a subset of the results obtained, specifically on pre-reduction tests done on South African chromite and manganese ore pellets.In this set of tests, it was found that the use of an organic binder, together with B2O3 as flux, promotes manganese ore pre-reduction, but retards the pre-reduction of chromite. By pre-reducing the manganese ore feed, significant cost savings can be achieved in subsequent smelting processes and there is potential to minimise manganese losses as a result of fuming, due to possible lower operating temperatures."

Jan 1, 2016

By W. R. McQuarry, M. E. Wilson

"The structural relations of the ore deposits of the Staclacona mine are of considerable interest, for they not only illustrate the effect of a change in competence of the wall-rock, but also the effect of either the convergence together of two faults, or a change in the clip of a fault, or both of these factors, on ore deposition.The mine is in claim M.L. 1818 near the eastern end of Pelletier lake and lies about 2 miles south of the town of Rouyn. The Company was organized in 1926 to explore claims in Rouyn, Dufresnoy, and Beauchastel townships, but attention was directed mainly to those in Rouyn. Underground work began in 1928 with the sinking of the main shaft to a depth of 300 feet. Except for 1930 and 1931, the mine has been in operation continuously since that time. Production, which began in 1936, up to the end of 1946 has amounted to 1,381,422 tons of ore from which gold and silver having a total value of $7,861,860 were recovered. The average grade of the ore milled in 1946 was $6.68.GEOLOGYThe rocks in which the Stadacona ore deposits occur are mainly volcanics of the (Keewatin) Abitibi series intruded here and there by numerous dykes, sills, or masses of diorite. Single dykes of olivine diabase and lamprophyre are also present in the property and are intersected in the mine workings. The volcanic rocks are part of a wide belt of northward dipping lavas and pyroclastic rocks, having a thickness of over 16,000 feet, that extend from the east-west trending belt of Timiskaming sediments on the south northward to the Horne Creek fault and form the south limb of a large syncline. The volcanic rocks within the property consist of numerous, east-west trending, northward dipping flows of andesite and a few flows of rhyolite interbedded with rhyolite ruff and agglomerate. The angle of dip throughout most of the property is 60° to 80°N., but in the main cross-cut on the 850- and 975-foot levels stratification in rhyolite tuff was noted dipping 35°N. The tuff at this point lies above an intrusive mass of diorite and it is probable that the change in dip results from the way in which the tuff and agglomerate have been thrust aside by the intrusive."

Jan 1, 1954

By R. Jacobs, R. E. L. Pobe

The Moab Khotsong Metallurgical Complex comprises a gold recovery circuit (Noligwa Gold Plant) and a uranium recovery circuit (South Uranium Plant). The metallurgical complex employs the reverse leach technology for the recovery of uranium; specifically, uranium extraction prior to gold using acid leaching processing. The inherent uranium species in the ore are uraninite, coffinite and brannerite. The slurry that has been subjected to the uranium leach is further subjected to solid-liquid separation to separate the uranium in solution through counter-current decantation and the solids (solids carrying all the gold), which are repulped and returned to the gold plant for gold processing. The uranium leaching process is critical to the uranium playing a pivotal role as a value adding by-product to the gold operation. The benefit of adopting this reverse leach process, for improved gold recovery, is that it exposes gold particles occluded within the uranium matrices and avails them for cyanidation. However, the drive for operational efficiency in uranium extraction must be weighed against the incremental operating cost against the backdrop of depressed spot uranium market prices. Over the years, efforts were directed at maintaining the benefit of extracting uranium to enhance gold recovery and simultaneously, optimizing the operating cost model for the uranium operation to remain relevant to the gold operation. The aim of this paper and the corresponding presentation is to give account of the uranium processing optimization journey embarked on over the years; namely, the operational trade-offs made in the process of ensuring that uranium continues to add value to the gold operation. These challenges were faced while striking a fine balance between processing cost and operational efficiencies against the prevailing depressed uranium spot market prices. Mineralogical- and laboratory tests conducted form the support basis for the results and observations articulated in the paper.

Jan 1, 2020

By Frédéric Flament

The aim of this paper is to present the development of a data validation software as component of a commercial mass balance package. Data reconciliation through mass balance is basically an optimization problem constrained by the law of mass conservation. As any optimization problem solved with a modern and robust software application, erroneous information or starting point will always lead to a solution, whether it is realistic or not. Moreover, good information wrongly weighted by an incorrect error model will lead to the same bad ending. In an industrial context where competition requires accurate information, data reconciliation without validation can be disastrous and calls for data validation before and after reconciliation. The challenge in automated data validation resides in the development of general theoretical solutions. Until now, none has been presented and only partial solutions are available. The development presented in this paper incorporates some of these partial solutions. The integration of human knowledge into the application paved the way to a more general solution. The combination ?theoretical validation ? human expertise? proved to be flexible and reliable enough to be integrated in a commercial data validation package. Numerous examples though industrial applications are presented. These applications demonstrate the validation of raw measurements, user support in the development of error models and validation of reconciled measurements.

May 1, 2001

By H. Link

"The extremely low permeability of the water-carrying beds in the nevonian formation of the potash district of Saskatchewan does not permit the injection of cement or chemicals to seal off sufficient water from the shafts to guarantee the safety of the mine. Based on many years of experience with wet shafts in the German and French potash mining industry, Alwinsal Potash of Canada Ltd. decided to have their potash mine shaft near Lanigan, Sask., which was completed in September of 1967, lined by means of an absolutely watertight lining of combined concrete and steel-plate construction.This shaft lining, consisting of two steel shells of welded construction with concrete filled in between, has proved itself before in the German coal mining industry. It guarantees a permanently dry shaft. With this type of lining, time-robbing and very often abortive injections are superfluous and the regular repair and adjusting work as well as the corrosion of conveyances and ropes can be avoided. Above all, however, the danger of a water inrush into the shaft is abolished completely."

Jan 1, 1971

By H. R. Huffman

"This paper reviews the history of iron powder production, and outlines the varions production methods. The uses of iron powder and the reasons for its use are also covered. The paper concludes with a look at future trends in powder metallurgy.INTRODUCTION TO POWDERSTHE IRON POWDER INDUSTRY is enjoying a very rapid growth. The annual consumption of iron powder is now at the 100,000-ton level, twenty times higher than it was at the end of World War II, and the yearly in-crease bas been almost consistent at 20 per cent. This is expected to increase, bringing the tonnage to 200,-000, or even (according to some experts) to 270,000, in 1970. In any event, powder metallurgy is expanding at high speed and making an important impact on the production of iron and steel products. Some indication of the faith in the future market may be indicated by the commencement of three new powder production plants in Canada with a total capacity of 136,000 tons annually."

Jan 1, 1969

By Holly L. O. Huyck

Disseminated hypogene sulphide mineralization in the Lakeshore porphyry copper deposit (also called the Cyprus-Casa Grande deposit) of Arizona occurs in both a granodiorite stock and associated volcanic rocks. Structural development and vein paragenesis at Lakeshore are similar to other Laramide porphyry copper deposits of the southwestern U.S. Two major structural trends are present: intrusive contacts, copper-bearing potassic alteration veins and major post-mineralization Basin-and-Range type faults trend N-NW, whereas phyllic veins and secondary post-mineralization faults trend N-NE, and deep propylitic veins trend NE-ENE. Lithologic permeabilities, caused by fracturing in all rocks and by grain size distributions in volcanic rocks, controlled the distribution of copper mineralization. Differences in chemical compositions of host rocks did not affect are grade. Structural control, through density of chalcopyrite-rich veins, was the major control of are grade. Fracture density of other potassic vein types did not strongly affect grade. Within volcanic lithologies, permeability, as indicated by grain size distribution, provided a second-order, but significant, control of mineralization.

Jan 1, 1990

By Jason Mewis

Jason Mewis is the founder and president of ENGCOMP, a consulting engineering firm in Saskatoon Saskatchewan focusing on the mining sector. He has over 17 years of experience working on projects for the heavy industrial market in Saskatchewan, as both a Senior Structural Engineer and Project Manager. Jason is an entrepreneur with experience in many areas including estimating, planning, facilitation, and Risk Analysis. He has presented at many seminars and conferences across North America on the subject of Project Risk and and has become recognized for his expertise in this area of project management.

May 1, 2013

By F. Habashi

"A short account will be presented on the extraction of rare earths from monazite sand, bastnasite ore, and phosphate rock of igneous origin. This includes mineral beneficiation, leaching methods, fractional crystallization [of historical interest], ion exchange, solvent extraction, precipitation from solution, and reduction to metals. INTRODUCTIONOriginally, the term rare earths was only used for the oxides, R2O3, which are similar to one other in their chemical and physical properties and are therefore difficult to separate. Within the rare earth group, the elements scandium, yttrium, and lanthanum differ in their atomic structure from the elements cerium to lutetium (the lanthanides, Ln). Scandium occupies a special position with respect to this classification and its other properties, and therefore does not belong to either of these groups. The rare earth elements always occur in nature in association with each other. The isolation of groups of rare earth elements or of individual elements requires costly separation and fractionation processes owing to the great similarity of the chemical and physical properties of their compounds, which explains why the history of their discovery has extended over such a long period.The word “rare”, when used to describe this group of elements, originates from the fact it was thought that these elements could only be isolated from very rare minerals. Considering their abundance in the Earth’s crust, the term rare is now inappropriate. These elements are lithophilic and are therefore concentrated in oxidic compounds such as carbonates, silicates, titano-tantalo-niobates, and phosphates.The abundance of the rare earth elements taken together is quite considerable. Cerium, the most common rare earth, is more abundant than cobalt. Yttrium is more abundant than lead, whereas Lu and Tm are as abundant as Sb, Hg, Bi, and Ag. For all practical purposes promethium does not occur in nature. It forms only in nuclear reactors."

Jan 1, 2012

By C. Edwards

"The processes for rare earth (RE) extraction include the decomposition of RE ore concentrate minerals, and the concurrent or subsequent leaching of the rare earth elements (REEs) from the decomposed minerals. Ore beneficiation concentrate can be decomposed by, for example, acid roasting, caustic cracking, and chlorination. The REEs can be selectively extracted. The nature of the REE extraction process depends on the concentrate mineralogy and the targeted products. Based on the experience of the Saskatchewan Research Council (SRC) in RE processing, this paper reviews REE extraction from the major RE minerals, including bastnaesite, monazite, xenotime, ion-adsorption clay, allanite, cerite, and eudialyte. An update on the development of RE extraction processes is also provided.INTRODUCTIONRE processing usually involves two operation stages, physical processing and chemical processing. Physical processing, also known as ore beneficiation, separates the RE minerals from gangue minerals to produce an RE mineral concentrate. Chemical processing, which includes extraction and separation processes, converts the RE mineral concentrate to an RE compound which is either an end product or an intermediate product for the subsequent production of individual REEs or other compounds. The RE extraction process uses one or more reagents to decompose the RE minerals and to leach the REEs into solution. The RE separation process uses solvent extraction, ion-exchange, or chemical precipitation to produce either rare earth oxides (REOs) or individual REEs. The authors have previously reviewed RE ore beneficiation processes (Zhang & Edwards, 2012). This paper focuses on extraction processes for a variety of RE minerals. Table 1 gives the chemical compositions of the RE minerals mentioned in this paper. The varieties of RE extraction methods use many different reagents. These reagents are mainly inorganic acids, alkalis, salts, and chlorine gas. The commonly used acids include sulfuric acid (H2SO4), hydrochloric acid (HCl), and nitric acid (HNO3). Sodium hydroxide (NaOH) and sodium carbonate (Na2CO3) are the most common alkalis. Salts include ammonium sulphate ((NH4)2SO4), ammonium chloride (NH4Cl), and sodium chloride (NaCl). HCl is normally used to extract RE from silicate ore minerals such as allanite and gadolinite. The alkalis and H2SO4 are mainly used to leach REs from phosphate ores minerals like monazite and xenotime. The carbonate ore RE minerals such as bastnaesite can be treated using either acids or alkalis. HNO3 is mainly used to leach eudialyte and apatite. The extraction of REs from ionadsorption clay deposits uses salt solutions. Chlorine gas (Cl2) is exclusively used in the RE chlorination process which can be used to treat any RE minerals."

Jan 1, 2012

By J. D. Clucas, D. B. Dreisinger

"The Starved Acid Leach Technology (SALT) has been developed by Search Minerals to extract nickel and cobalt VALUES (from nickel laterite ores (saprolites) and Caron plant residues. The technology consists of sulfuric acid atmospheric leaching of the ores or residues followed by neutralization and recovery of nickel and cobalt. The acid addition is carefully controlled to maximize nickel and cobalt extraction while minimizing overall consumption of acid by various gangue reactions. The final recovery processes can be tailored for maximum economic benefit. The use of resin-in-pulp processing is advocated as the most cost effective method for initial recovery of nickel and cobalt from lower grade materials. INTRODUCTIONCommercial processes for laterite ore treatment vary according to laterite type and local conditions. The two general categories of laterite ore are saprolites and limonites.Saprolites are generally treated by pyrometallurgical processes for reduction of nickel to matte or ferro nickel alloys. The high cost of drying, chemical reduction and electric furnace smelting requires that the saprolite grade for economic processing is of the order of 1.8% nickel or higher. There are of course large amounts of saprolite that fall below this “cut-off” grade that are not currently economic.Limonites can be treated by the Caron Process which involves drying of the ore, a reduction roast of the whole ore to reduce nickel and cobalt (and some iron) to the metallic state followed by ammonium – ammonium carbonate solution to leach nickel and cobalt. Nickel and cobalt recovery are low (70-85% for Ni and 20-50% for Co). The Caron plant tailings therefore contain high levels of unrecovered nickel and cobalt. Relatively few plants have used the Caron Process and new installations are not anticipated due to the high energy cost of drying and reduction roasting. Nevertheless, plants in Brazil (Niquelandia), Australia (Yabulu) and Cuba continue to operate.The high pressure acid leach (HPAL) process was pioneered at Moa Bay in Cuba in the 1960’s in the treatment of limonites. In the 1990’s the HPAL process was touted to revolutionize the nickel industry with low capital and operating costs for new installations. Murrin Murrin, Cawse and Bulong in Australia were started and in each case the technical difficulties of operating at extreme temperature and pressure conditions and the challenges in “ramping up” a complex hydrometallurgical facility conspired to cause these plants to fail economically. At present, Cawse and Bulong are closed and Murrin Murrin continues to operate after undergoing a financial restructuring."

Jan 1, 2012

By A. Anderko, R. D. Springer, R. D. Young

A thermodynamic model has been developed for calculating phase equilibria and other properties of multicomponent electrolyte systems. The model has been designed to reproduce the properties of both aqueous and mixed-solvent electrolyte systems ranging from infinite dilution to solid saturation or pure solute limit. The model incorporates formulations for the excess Gibbs energy and standard-state properties coupled with an algorithm for detailed speciation calculations. The excess Gibbs energy model consists of a long-range interaction contribution represented by the Pitzer-Debye-Hiickel expression, a second virial coefficient-type term for specific ionic interactions and a short-range interaction term expressed by the UNIQUAC equation. The accuracy of the model has been demonstrated for common acids and bases and for multicomponent systems containing aluminum species in various environments.

Jan 1, 2005

By D. R. Boyle

"The basal type uranium deposits in south central British Columbia occur within unconsolidated, late Miocene fluvial paleochannel sediments that overlie major fault zones within the Okanagan Highlands Intrusive Complex. Prior to the formation of mineralization, paleovalleys containing host sediments were flooded by Pliocene valley basalts of the Plateau Basalt Formation. Five uranium deposits have been outlined to date, of which the Blizzard (4020 tonnes U) and TYee (650 tonnes U) are the largest. The basement intrusive complex underlying the deposits varies in age from early Cretaceous to Eocene and is comprised of quartz monzonite, granodiorite, Coryell monzonite, porphyritic granite and pegmatite.The mineralized host sediments comprise a complex sequence of interfingering conglomerates, arkosic sandstones and mudstones. Organic material is abundant in many, but not all, mineralized beds. Uranium mineralization is present in the form of uranous (ningyoite) or uranyl (saleeite, autunite) phosphates coating clastic grains and filling voids. Because of very strong reducing conditions related to large concentrations of marcasite and organic material, ningyoite is the only uranium mineral in the Tyee deposit, whereas the Blizzard deposit contains a more complex assemblage of minerals (saleeite, autunite, ningyoite). The observed paragenetic sequence of mineral precipitation in the Blizzard deposit (autunite-saleeite-ningyoite) indicates that the uranyl minerals, saleeite and autunite, are primary. Small amounts of Zn, Ni, Co and Mo have been introduced into the Blizzard deposit whereas Mo is the only non ore-forming element enriched in the Tyee deposit. Mn is the only depleted element in both deposits."

Jan 1, 1986

By A. Brown

"The first printed report on Canada's coal resources was published in Paris in 1672(1. Written by Nicholas Denys, then Governor for France of Isle Royale (Cape Breton Island), the report dealt exclusively with the coal of the Atlantic region because little or nothing was known about the interior of North America. Clearly, also, he wrote about a resource of which much was known among the military and early settlers of that day. This early knowledge is not surprising because the presence of coal is clearly displayed in the coastal cliffs of Cape Breton as it was also, at a later date, to the first arrivals on Vancouver Island.An example of early use was that of the French garrison at Fort Louisburg in the first half of the 1700s. Still earlier, in 1643, Governor Winthrop of New England reported the receipt of a small ship load of coal from New Brunswick (2). Despite this early knowledge there was no significant development of coal mines until the 1800s. This had to await the ending of wars between the French and British factions and the growth of population. Basically, also, its development awaited the beginning of the Industrial Age. Organized coal mining began in Canada in 1825 with the formation of the General Mining Association. This British firm was the creation of a group of London jewellers which had been granted royal rights, by the British crown, to all minerals in Acadia (substantially the present Maritime Provinces). The Association took over a number of existing small mines and opened up larger operations in the Sydney coal field. Subsequently, it expanded to the mainland Nova Scotia fields at Stellarton and Springhill. The Association continued in existence as the major coal producer in Canada until 1899 when it sold its properties to the Nova Scotia Steel and Coal Company, Limited (3). This Montreal-based firm managed the coal mines as a major Canadian producer of coal until 1968 when financial problems led to the expropriation of the mines by the Government of Canada. The federal government established the Cape Breton Development Corporation to operate the larger mines on Cape Breton Island. This Crown corporation continues to be a .major producer of coal in Canada."

Jan 1, 1985

By Pierre Sauvé, Guy Perrault, Armand Savoie, Pierre Trudel

"La Mine Doyon est une entreprise détenue à parts égales par Cambior Inc. et par Minerais Lac Ltée qui en est l'opérateur. Les statistiques de production au 1"" janvier 1990 indiquent que, depuis le début de son exploitation (1980), 10.3 Mt de minerai d'une teneur moyenne de 5,9 g Au/t ont été usinées. Les réserves sont estimées a 10,4 Mt à une teneur moyenne de 6,6 g Au/t jusqu' à une profondeur de-450m.Trois zones minéralisées sont présentement en exploitation. Ces zones sont caractérisées par la présence de veines de quartz-pyrite accompagnées localement de chalcopyrite. Les roches encaissantes sont les schistes a séricite (zone# 1), les roches volcaniclastiques de composition felsique (zone #2) et intermédiaire (zone #2, zone ouest) ainsi que les roches intrusives appartenant au pluton de Mooshla (zone ouest). Toutes ces roches ont subi un métamorphisme régional dont l'intensité correspond à la limite supérieure du facies schiste vert ( -425°C}.Les zones minéralisées sont situées à I' intérieur des zones de déformation et d'altération les plus intenses. L'étude des structures mineures nous indique que la région a subi plusieurs épisodes de déformation ductile a fragile qui sont antérieurs, contemporains et postérieurs al a minéralisation. L'analyse structurale indique également que la déformation majeure est associée à une importante phase de compression orientée N-S a NNE-SSW.L’ensemble des observations suggère que la minéralisation et I' altération associée sont de type épigénétique. La mise en place du minerai se serait effectuée à la fin de l'épisode principal de métamorphisme et de la déformation."

Jan 1, 1990