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By P. B. Marchant

"Considerable gold and silver values are lost to flotation tailing from both the South ern Tail orebody and the Main Zone orebody,The underlying mineralogical assemblage of the two ore types is quite different. This paper describes the research strategy employed to determine an economical method to scavenge gold and silver from flotation tailing.Much of the unrecovered gold in the Southern Tail ore is associated with arsenopyrite while silver losses are probably as unliberated tetrahedrite and minor silver-bearing sulphosalts. Processing alternatives tested include bulk sulphide recovery by flotation and gravity, selective pyrite-arsenopyrite separation, cyanidation, and oxidation pretreatments such as pressure leaching and bioleaching to enhance cyanidation response. Gold losses in the Main Zone ore are more closely associated with silver-bearing minerals, pyrite, and siliceous gangue and consequently respond more readily to conventional tank cyanidation. The 5 300 tid carbon-in-leach circuit to treat Main Zone flotation tailing is described in detail.IntroductionEquity Silver Mines Limited's open-pit silver mine, a 70% owned subsidiary of Placer Development Limited, is located 37 km southeast of Houston in north central British Columbia. The deposit consists of two major ore zones, Main Zone and Southern Tail (Fig. I). These zones are within a sequence of Cretaceous volcanic and sedimentary rocks striking 015 degree s and dipping 45 degrees west. Feasibility ore reserves were 28 million tonnes grading 106 g A g/t, 0.96 g Au !t and 0.38% Cu a strip ratio of 2.2 to I. A detailed mineralogical description has been presented and a comparison of the two ore zones is shown in Table I."

Jan 1, 1986

By W. P. Clement, M. C. Bell

"PT International Nickel Indonesia, a subsidiary of Inco Limited, mines and processes nickel laterite ores for the supply of nickel matte, primarily to the Japanese market. It has large ore reserves and a substantial portion of its energy requirement is supplied from its own hydro-electric plant. Since the early 1980s, the company has increased its production level and has dramatically reduced its unit costs. Projects are in progress to further decrease costs and recently a major expansion - 31% increase in nominal capacity - has been undertaken, consistent with marketing opportunity in Japan. The expansion project, which has a low capital cost per unit of production, is outlined in this paper.IntroductionPT Inco is a fully integrated mining and smelting facility located in Indonesia on the island of Sulawesi, and is based on a very large, world class nickel laterite orebody located 60 km inland from the Gulf of Bone. Construction of a single furnace line plant started in 1973. Complete supporting infrastructure of townsite, hospital, schools, shopping centres, road systems, and an airport, all had to be built. Prior to completing the first phase, the decision to expand the plant to a three kiln-furnace line was made. A major factor in the decision was the ability to develop the hydroelectric potential of the nearby Larona River. This involved construction of a 30 m high dam, and a 7 km surface canal to the generating station. The penstocks drop 150m to three 55 Mw turbines. The power generated is sufficient to supply the entire electrical load for the complete smelting complex and town site. The unique capability to supply the high smelting energy requirements for a lateritic nickel ore by fully captive hydro power, has, over the past decade insulated PT Inco from the large upward swings in oil prices suffered by other lateritic nickel producers.Wet ore is first partially dried to 20% moisture in two co-currently fired rotary driers, screened on an internal trommel, and delivered to two dry ore storage buildings. The partially dried ore is calcined at 750°C and partially reduced in three large 100 m by 5.5 m kilns. Reduction is accomplished by lancing bunker C oil onto the ore bed at the discharge end of the kiln. The kilns are equipped with five air pipes with onboard blowers to gradually bum the combustible gases generated by the lance oil. The kilns were also originally equipped with two side burners to supply additional heat close to the feed end of the kiln. Calcined ore is transferred by 20 ton refractory lined containers to the furnace feed bins at the top of the smelter. Smelting is carried out in three 18 m diameter furnaces powered by 45 Mw transformers. Slag is skimmed at 1550°C through water cooled breast plates and removed to the slag dump by Haulmaster Carriers. Matte of 25% Ni, 1O% S, 65% Fe is tapped at 1350°C into the converter aisle. The iron is slagged in converters to produce a nickel sulphide matte at 80% Ni. Three rotary Kaldo converters were initially installed. The product matte is granulated, dried, and shipped in returnable rubber bags to Japan. Most of the matte is fluid-bed roasted to nickel oxide and then directly reduced in a second reducing fluid bed reactor at our facilities in Japan.The smelting complex was designed to smelt two million tonnes of dry ore per year, equivalent to 34 000 tonne nickel production at an ore grade of 2% Ni. The cost of the entire project including working capital was $900 million."

Jan 1, 1990

By M. A. Mahtab, A. S. Burgess, R. G. Charlwood, P. O. Sandstrom, J. L. Ratigan

"Addressing a current operating and environmental concern, this paper first presents an overview of the design program for an underground radioactive waste storage vault. It then reviews the construction, operational and thermal-mechanical aspects of such a project. The relevant experience from major underground construction projects is reviewed and the paper concludes with brief comments on the design-oriented research and development requirements . 1. IntroductionA program to develop techniques for the ultimate disposal of high-level radioactive waste into geologic formations was announced by AECL in 1975 (Tamrnernagi, 1976), and is discussed more fully in a later document (Boulton, 1978). Our paper summarizes the studies caried out, as part of the AECL waste management program, by the authors and associates on the development of design concepts for a vault situated in crystalline rocks of the type present in the Canadian Shield.The design studies have been carried out in three phases, the elements of which are described in Section 2.1 (also see Fig. 1). The general vault designs are presented in Section 2.2. These provide for emplacement of either immobilized fuel or solidified reprocessing wastes in a room-and-pillar facility at a depth of 1,000 m.Section 3 reviews the principal design considerations from the construction, operation and thermal-mechanical response aspects. The concepts have been studied in both temporal and spatial domains, and the requirements and responses assessed in each as appropriate.Section 4 briefly reviews the availability of relevant experience from previous major underground construction projects. Finally, in Section 5, brief comments are given regarding design-oriented research and development requirements."

Jan 1, 1980

By G. Ladouceur

"Northgate Patino Mines produces a copper concentrate assaying 22% Cu. Moisture content of the product averages 12.5% H20 when dewatering with conventional vacuum disk filters. Recent smelter contracts reflect higher energy costs and increased moisutre penalty charges. This situation forced Northgate to investigate possible alternatives such as rotary dryers and pressure filtration. After laboratory testing, it was found that pressure filtration could economically meet the smelter's requirements. Pilot plant testing was then undertaken and confirmed the laboratory test results. Based on these findings, a model I.S.F. 1500 x 28/22 I.R. Lasta filter press was purchased.This paper describes the Copper Rand dewatering flowsheet the operation of the press and the economics of the process.IntroductionNorthgate Patino Inc. operates 2 und erground copper mines and a 3300 ton per day concentrator located in Chibougamau, Quebec, some 425 miles north of Montreal and 300 miles northeast of Noranda, Quebec.Copper concentrate is produced at a rate of 250 tons per day and trucked to a C.N. siding 10 miles from the mill site and then railed to the Noranda Smelter.Until early 1981, moisture content of the concentrate at 12.5070 was acceptable since smelter penalty charges were relatively low ($0.10 per ton per unit over 10070 moisutre). In January 1981, a new contract was signed in which the penalty was raised to $1.10 per ton per unit over 8070 . This and the increasing truck and rail costs justified the change-over from vacuum to pressure filtration . The Lasta I.S .F. 1500 x 28/22 was then purchased and installed, and operation of the plant started in December 1981."

Jan 1, 1987

By D. MacGREGOR, VERA TUREK

The Badgeley Island quartzite deposir in Georgian Bay is worked by lndusmin Limired and was developed in 1969 to produce a high-qualiry silica for use in rhe ferrosilicon and glass industries. The quartzire is part of the Precambrian Bar River Formarion, and is a generally whire, fine-grained rock containing rhin sericire seams. Primary sedimenrary features wirhin the deposit are bedding and ripple marks. Amphibolite and diabase dykes cur the quarr zire body.

Jan 1, 1984

By M. K. Aly

"ABSTRACTThis paper briefly describes the mining method proposed for the iron ore deposits in Wadi Shatti, western Libya. The iron ore produced will be used in the steel plant at Misurata on the Mediterranean coast. GeologyThe Wadi Shatti is located in the central part of western Libya. Flat iron-bearing sedimentary rocks are exposed with remarkable regularity in this valley. It occupies an area 200 km long by approximately 20 km wide trending in an eastnortheast direction between Jabal Fazzan in the north and Zel1af Sand Dunes in the south (Fig. I). Only a part of this valley has been ful1y explored by drilling on a 400- by 200-m pattern in an area of 18 km2 south of the town of Taroot. Farther to the south, no exploration by drilling can be done due to the presence of sand dunes .The iron-bearing bed is about 10to 15m thick, with a dip of 2 to 4 degrees in a southerly direction . The overburden consists mainly of clayey sandstones, 4 to 5 m thick, in the northern part, increasing to 35 m in the southern part. The iron ore bed consists of three different layers, from the floor upward: (1) Lower Reduced Ore (LRO), having a varying thickness of 1 to 7 m; (2) Magnetic Ore (MO), with a thickness varying from 4 to 12 m mainly in the central part of the area; and(3) Upper Reduced Ore (URO), with a bed thickness varying from 0 to 7 m."

Jan 1, 1980

By M. P. Brameld, G. Archambault, R. H. McCorkell

"The rates. J. at which materials emit radon are of interest in studies of the sources of radon in mines. buildings. soils overburden and other places where it may indicate the location of ore or be a health hazard. Here J values were determined by measuring the radon concentrations by means of scintillation counters or alpha-particle track detectors in chambers in which specimens were sealed. Errors in the measurements due to leakage from the chambers. lack of secular equilibrium in the decay series of the radon measured, or the limited rate of diffusion of radon from powders or within the chambers were shown to be insignificant or could be corrected. Scintillation counter measurements are highly reproducible; track detector measurements also appear useable and may be made on many small samples using small, simple chambers. J measurements were made for uranium ores, drill cores, standard CANMET pulps. concrete and gypsum board. Powdering of ores had little effect on their J values.IntroductionIn the mining and mineral sciences, radon (222Rn) is of interest chiefly for two reasons: in geochemical exploration for mineral deposits its presence in soil. overburden and water in more than background concentrations may indicate the existence and locations of uranium deposits and, if present in sufficient concentrations, it is a possible health hazard in mines, mills and other places where people live and work. The latter aspect of radon's role has received considerable attention recently(1,2,3,4,5), and surveys of soil-gas radon concentrations have been used for some time in searches for uranium deposits(6,7,8,9,)."

Jan 1, 1981

By D. R. Murray

"A promising long-term solution to uranium tailings reclamation has been the establishment of vegetation directly on the waste material. The soil profiles of areas vegetated for 0, 3, 8 and 10years are characterized as to the extent of soil development . The characteristics are compared with the visually interpreted soil profile. Organic matter content is presently the best analytical judge of soil development. Vegetation procedures have produced a build-up of organic matter in the top 10cm of tailings where natural microbial populations flourish. The CN ratio is high and the profile depth shallow, suggesting the need for additional fertilizer applications.IntroductionThe reduction of surface erosion and improved aesthetics of uranium tailings has been accomplished by the establishment of a vegetative cover. In 1971 research concentrated on developing methods of plant establishment at minimal cost and soil amendments on uranium tailings in Elliot Lake , Ontario. One process of stabilization used on a large scale was direct seeding of grasses and legumes into the tailings which had received only limestone and fertilizer treatments. Fertilizer additions were made for varying lengths of time in subsequent years to maintain the established cover.Maintenance-free vegetation would be possible once a soil profile developed with sufficient nutrient recycle and plant regeneration capabilities(1). Tailings vegetated in 1971, 1973 and 1978 that had received either 3 or 5 years of maintenance fertilizers were examined for soil development profiles. The extent of soil development and the relative effect of time and fertilizer would provide direction on reaching the goal of a self-sustaining cover.This paper is an abridgment of Division Report MRP /MRL 81-126 of the same title(2)."

Jan 1, 1984

By J. E. Armstrong

"Pinchi Lake and Bralorne Takla mercury mines; Canada's only producing mercury mines during the past war (1939-1945); occur within the Pinchi Lake mercury belt. This belt is in central British Columbia; and is approximately 150 miles long extending from southeast of Fort St. James northwest to the headwaters of Omineca river.During the field season of 1937 cinnabar was discovered by J. G. Gray of the Geological Survey in Permian limestone on the north shore of Pinchi lake, where the main showings of the Pinchi Lake mercury mine were later developed. Following the publication of Gray's report claims were staked in the spring of 1938, and later that year were optioned to the Consolidated Mining and Smelting Company, and in June 1940 a reduction plant with a rated capacity of SO tons a day was brought into operation. In 1943 the plant had a capacity of more than 1,000 tons a day. The grade of ore treated was reported to average S to 10 pounds of mercury a ton, and production during the years 1940 to 1944 exceeded 4,000,000 pounds. The Bralorne Takla mercury mine, which is 90 miles northwest of the Pinchi Lake mercury mine, was brought into production in November 1943 with completion of a 50-ton mill. During the next 10 months more than 100,000 pounds of mercury were recovered, and the grade of the ore was reported to average 15 to 20 pounds of mercury a ton. During the war, production from these two properties was far in excess of Canadian requirements, and Canada was able to supply the United Kingdom and United States with part of their need.s. Both properties ceased operations during 1944, and, due to an excessive supply of mercury, have not yet reopened. The Pinchi Lake mercury belt lies along a northwest-trending major fault zone called the Pinchi fault zone (Fig. 1). This zone is approximately 150 miles long and varies in width from 200 to 1,000 feet . It forms the northeastern edge of a northwest trending fault belt up to 10 miles or more wide in places. In general the northeastern margin of the Pinchi fault zone represents the contact between closely folded, stratified Permian rock:s on the southwest and Mesozoic formations and Omineca intrusions on the northeast. It seems probable that the fault zone marks the site of major thrust faulting from the southwest and that the Permian rocks moved up with respect to the Mesozoic formations. Intense faulting occurs in the Permian rocks within the fault zone. There the more important faults trend north and northwest, dip steeply west and southwest, and may join a major low-angle thrust fault at depth."

Jan 1, 1954

By S. A. Willis

For over 150 years gypsum from Nova Scotia has been shipped via water to the United States east coast. The history of the movement has been one of continual improvements in quarrying, loading, and shipping. At the start, man-size rock quarried by hand auger and black powder was loaded into horse-drawn carts, hauled to the waterfront, and dumped into 100-ton sailing schooners. Last year the Canadian Gypsum Company completed the latest and most modern of these successive installations for the movement of this Nova Scotia product. The new system loads crushed gypsum into 11,600-ton vessels at Hantsport in two hours' time at high tide. At the head of the bay of Fundy are located vast deposits of gypsum. The tidal rise and fall of 45 feet twice a day, with its attendant swift currents, has always presented major navigation and shipping problems. When the tide is low, the Hantsport ship-loading pier is high and dry, with no water within 400 feet. Three hours later a large river has flooded in (running up hill) and large steamers can dock at the wharf. Before high water, the ship has been loaded and -there is a depth of 30 feet or more of water around the wharf. At high water, the ship sails away on the ebbing tide as the river runs in the other direction. Tides are caused by the gravitational attraction of the moon and sun on the ocean. When both celestial bodies are pulling together the 'spring tide' results, which means a 50-foot variation from high to low. This occurs at new and full moon. A week later, when the sun and moon work in opposition, the difference is only 32 feet between high and low water. The bay of Fundy, having a funnel shape (150 miles wide at mouth from Yarmouth to Portland, while the upper estuary at Hantsport is only one mile wide), builds up a tidal influence from 10 feet at the mouth to an average of 45 feet at Hantsport.

Jan 1, 1948

By S. A. Willis

For the past several months 1 have been considering the functions of this Society and its reasons for continuance. The fact that it is now sixty years old indicates that it must have been of value and we all know that it has contributed much that was worthwhile to Nova Scotia and the Maritime Provinces. The question arises: "What is the future of the Society, and how might its activities be altered to suit present-day activities and to be of even greater value?" Let us analyse the various activities in which we are interested and determine which, if any, are of more value than others. First of all, at least once a year we have a two-day convention which constitutes a holiday for the members and their wives. This has always been well planned and has been much enjoyed. Second, there are many social and business contacts made during the meetings and interim meetings. These certainly tend toward better relations and co-operation between the various people concerned. Mine operators meet other mine operators, engineers, people from the education fields, Department of Mines, and so forth. Various sales contacts are made at these meetings which can be of value and use to both the Company selling and the customer. Third, and of greatest importance in any conference, is the exchange of ideas by the members and by authorities from outside, and the presentation of slides and drawings, or actual visits to operations. It is upon this phase that 1 wish to elaborate. These papers are not only presented to our meetings but are generally published in the C.I.M. Bulletin so that members throughout the Dominion have access to them. Many important papers describing new phases of operations have originated through our Society.

Jan 1, 1947

By Fernand Dubuc

"The New Pascalis property is situated in Louvicourt and Pascalis townships, about 200 km east of Val d'Or, Quebec. It is made up of 38 claims under option to SOQUEM since August I, 1981 during which time SOQUEM has earned a 60% interest. The property is underlain by archean basaltic volcanics that have been intruded along the west boundary by the Bourlamaque batholith, and by north -trending dykes.The 1982 disco very is related to the re-evaluation of a 1931-gold showing located in the southeastern part of the property, a combination of the presence of a weak IP anomaly at the eastern end of the showing and overburden stripping at that location, resulted in the discovery of a mineralized diorite dyke.The dyke, which strikes N 20 ° Wand dips vertically , has intruded E-W striking agglomerates. The dyke contains south dipping, gold-hearing quart z-tourmaline-pyrite veins that are distributed in a ladder-type arrangement. Visible gold may he observed in drill core, and is generally associated with pyrite porphyroblasts.Diamond drilling has outlined th e dyke for a strike length of 300 11/ and a width of about 10 m, Result s of this drilling indicate, to a depth of 155 m, reserves of 834 000 metric tonnes grading 8.25 g Au/t un cut or 5.38 g Au/t cut to on e ounce. However, four deeper holes have added depth extension to 400m. Also, west of the dyke and within 150 m, other promising gold values have been intersected within quartz-tourmaline veins associated with other parallel dykes."

Jan 1, 1985

By W. E. Cockfield

"The Stump Lake mining camp is in Nicola mining division of southern British Columbia. It lies about halfway between the cities of Kamloops and Merritt.The deposits at Stump Lake are among the earliest lode discoveries of British Columbia, many of the claims having been staked between 1882 and 1885. A considerable amount of work was done prior to 189Cl, after which very little was done until 1916. Since that time mining has been carried on at intervals, much of it on the same deposits, by a number of companies, including the Donahue Mines Corporation, the Planet Mines and Reduction Company, the Nicola Mines and Metals Company, the Consolidated Nicola Goldfields, Limited, and others. According to the British Columbia Department of Mines, total production from the area has been: gold, 8,520 oz.; silver, 259,413 oz.; copper, 40,822 lb.; lead, 2,208,859 lb.; and zinc, 367,869 lb. Most of the production has come from the Enterprise vein and a branch of the Enterprrsethe King William vein."

Jan 1, 1954

By B. Van den Bogaert, K. Binnemans, T. Van Gerven

Europium is one of the most critical rare-earth elements due to the limited supply of this element combined with a high demand for red lamp phosphors. Therefore, the recycling of europium from end-oflife lamp phosphors has become increasingly important. Photochemical reduction of europium(III) to europium(II) and subsequent removal from rare-earth mixtures by precipitation as EuSO4 is a very selective and efficient method to recover europium from binary europium/yttrium solutions. It is shown that full separation of europium and yttrium is feasible from laboratory prepared non-equimolar Eu/Y mixtures up to a ratio of 1/20, which reflect the composition of red lamp phosphors. Optimal conditions were determined at pH 4 with isopropanol as a radical scavenger, which is a less harmful and toxic compound than the commonly used formic acid scavenger. Europium removal efficiencies of more than 95% were achieved, with a purity of 98.5% of the EuSO4 precipitate. The irradiance of the used light source, expressed in mW/cm2, has a large impact on the illumination time needed to remove europium from the solution. It is observed that higher irradiance results in a shorter induction time, i.e. a period during which virtually no europium is removed, and a faster removal rate once the induction time has passed. The photochemical separation technique was applied to recover europium from leached industrial red lamp phosphor (YOX) powder, with a Eu/Y ratio of 1/30. Europium recovery of 50% was reached with a purity of 96%, proving the feasibility of this method for the recycling of europium from end-of-life lamp phosphors.

Jan 1, 2016

By L. Avramovic, B. Han, R. Jonovic, Z. Stevanovic, K. Haga, S. Dimitrijevic, A. Shibayama, L. Godirilwe, Y. Takasaki

Pyrite (FeS2) and chalcopyrite (CuFeS2) which could not be recovered, are present in mine tailings discharged from Bor copper mine. About 0.3% of copper (Cu) is contained in these tailings, therefore it can be considered as a useful copper resource. At the same time, pyrite is known to promote acid mine drainage, therefore it is necessary to detoxify the tailings. An efficient method has been proposed to recover copper and reduce the environmental impacts of tailings by a combined hydrometallurgical process of high pressure leaching (HPL), solvent extraction (SX) and electrowinning (EW) process. HPL process yielded high copper leaching rate (>95%) when water (H2O) was used as a leaching media. On the other hand, pyrite contained in the tailings was converted to iron oxide by high pressure leaching. Stability evaluation of the residue obtained from high pressure leaching, confirmed almost no elution of metal ions. The percentage of copper extracted and stripped from the Cu-loaded organic phase reached 97.4%. Moreover, it was confirmed that a low iron concentration in the electrolyte of copper electrowinning, resulted in the improvement of current efficiency of copper electrodeposition. INTRODUCTION Mine tailings are a by-product obtained from the flotation process and there are billions of tons of already existing tailings all over the world. Moreover, several billion tons of additional copper tailings will inevitably be produced in the coming years (Onuaguluchi & Eren, 2012). With the declining average copper ore grade, alternative sources of copper are highly sought, mine tailings have a comparable copper grade to that of low-grade ores (0.2–0.3%) therefore, it can be said that copper mine tailings present a very important source of secondary raw material for copper production (Antonijevi´c et. al., 2008). Consequently, it is necessary to confirm the possibility of the valuable metals (such as copper) recovery from mine tailing in which copper content is about 0.2–0.3%. Our group investigates the flotation and high pressure leaching processes for recovering of copper from mine tailings. Previous work on related samples include copper and nickel flotation tailings from Botswana, and a low–grade copper ore sample from Chile. Impurities, which are associated metals in mine tailing such as iron (Fe) and aluminum (Al) co-dissolve into the solution during the HPL treatment process. Therefore, separation of copper from impurities is necessary for selective extraction of copper from mine tailings and upgrading the copper concentration in the leachate to enable further purifying of Cu by electrowinning. Thus, it is important to investigate the possibility of applying solvent extraction for concentration of copper from leachate solutions of mine tailings (Onuaguluchi & Eren, 2012; Antonijevi´c et. al., 2008). In this study, LIX-84I was used as an extractant of copper from the leachate of high pressure leaching of mine tailing. As a fundamental solvent extraction investigation, the extraction behavior of copper was studied using a simulated solution of the leachate from the high-pressure oxidative leaching of mine tailing. The effects of; solution pH, extraction time, extractant concentration and phase ratio on copper extraction and sulfuric acid (H2SO4) concentration on stripping were investigated.

Jan 1, 2019

By T. R. Yu, D. B. Counter

"Kidd Creek Mines primarily employs sublevel blasthole stoping for underground mining operations. During pillar recovery in No.1 Mine and panel mining in No.2 Mine, the free standing fill exposures may reach a height of 100 m or more. Consequently, a competent consolidated rockfill has been required to fill the mined openings.Since 1976, a total of more than fifty stopes and pillars have been filled with over four million tonnes of backfill. Fill exposures, up to heights of 100 m, have stood up well after blasting adjacent to the backfill.The fill material consists of a graded aggregate, Portland cement and ground slag. Sand has been occasionally mixed in the fill material to serve as avoid filler. The aggregate is prepared in a fill crushing plant, transported underground through raises, and hauled to the filling stopes by conveyors in No. 1 Mine and by trucks in No.2 Mine.Two mixing methods are currently in use. No. I Mine utilizes the free fall of gravity; the fill materials are sprayed with cement slurry as the aggregate departs from the conveyor into a mixing culvert. The other is the use of a rotary mixer on a batch basis in No. 2 Mine.This paper describes the fill systems, filling operation s, and the backfill quality control practice, together with various testing results related to improving fill quality as well as reducing fill costs.IntroductionThe Kidd Creek minesite, located 27 km north o f the city centre of Timmins, Ontario, produces 4.5 million tonnes of copper, zinc, lead, and silver ores annually from its No. 1 and No.2 underground mining operations. Both mines employ a sublevel blasthole stoping method."

Jan 1, 1983

By M. Honarmand, Z. Moezifar, H. Ranjbar

Abstract - The Sar Cheshmeh area covers an area of 640 km2, and is located in the southern part of the Iranian Volcanic-Sedimentary Belt. ETM+ images have been used for alteration mapping. The Crosta method was found useful for enhancing exposed areas containing hydroxyl and iron oxide minerals. However, field checking has shown that this method is not able to identify all the known areas of hydrothermal alteration. Many of these altered areas are also not clearly resolved in the geophysical data. In order to improve resolution of such areas, integration of remote sensing and geophysical data may be helpful. The Darrehzar Cu deposit was chosen as a control area. The data have been divided into two sets: explanatory and target variables. The explanatory data are Landsat hydroxyl and iron oxide images and airborne geophysical data (K, U, and Th total counts, and total magnetics). The target variables are rock geochemistry and ground geophysics. Data analysis and integration of geophysical and remote sensing data were performed using directed principal components. This technique is found to be useful for the delineation of hydrothermally altered areas with more confidence.

Jan 1, 2002

By S. M. Bulataovic, D. M. Wyslouzil

"The Martison Lake phosphate-niobium deposit is composed of unconsolidated residuum containing clay and silt-size particles of phosphate, niobium, iron and clay minerals. This paper deals with the laboratory and pilot plant development of the flowsheet and the reagent scheme to treat the Martison Lake ore. The ore, in addition to the complex gangue composition, contained several phosphate minerals including fluorapatite, francolite and florencite. The detailed laboratory and pilot plant evaluation resulted in the development of an effective flowsheet and reagent scheme for the selective flotation of phosphate minerals using a fatty acid circuit, followed by niobium flotation using amine collectors at acid pH. The developed flowsheet included des liming of ore, magnetic separation, and selective P 2 O 5 -Nb 2Os flotation. The main aspects discussed include characterization of the ore, step by step development of flowsheet and evaluation of dfferent modifiers and collectors for P20s and Nb20s flotation.INTRODUCTIONMartison Lake phosphate-niobium ore deposit, located 75 km north of Hearst, Ontario was discovered in the early 70's and has been the subject of extensive technical-economic evaluation in the late 70's and early 80's.The deposit is covered completely by 30 m to 90 m of glacial overburden and is composed entirely of unconso1idated material called residuum, that lies on top of a precambrian age carbonatite intrusive from which the residuum was derived by weathering and natural leaching in situ. The residuum is and irregularly-shaped horizontal lens, 2 km by 3 km in area, and up to 100 m thick, composed mostly of unconsolidated clay and silt-size particles of phosphate, niobium, iron, mica and clay minerals."

Jan 1, 1986

By P. R. A. Andrews, R. K. Collings

"In cooperation with le Ministère de l'Énergie et des ressources, Québec, le Centre de Recherches minérales, Sainte-Foy, Québec, and various silica producers, selected deposits of Quebec silica were identified in the southern region of the province and sampled to provide material for evaluation as glass and foundry sand by CANMET. The samples included both sandstone and quartz/quartzite. Head sample analyses ranged from 94.66% to 99.54% SiO2, 0.038% to 0.230% Fe2O3, and 0.16% to 4.51% Al2O3.Impurities included pyrite, iron oxides, and iron staining on quartz grains; carbonates as calcite; and alumino-silicates as clay and/or feldspar. Processing included size reduction to grain size, about 600 µm, by jaw, roll and cone crushers, and grinding in a muller-mixer. Removal of fines was achieved by attrition scrubbing and washing. Further removal of impurities was achieved by reverse flotation of iron oxide and alumino-silicate minerals employing both anionic and cationic collectors. Flotation resulted in a substantial improvement in the quality of most sands, the quartz concentrates analyzing 99.08% to 99.80% SiO2, 0.014% to 0.048% Fe2O3, and 0.06% to 0.49% Al2O3. Removal of iron oxide minerals was also investigated using the Sala HGMS. Iron oxide analyses of the non-magnetic fractions were comparable to the iron oxide analyses of the quartz concentrates obtained by flotation, 0.013% to 0.054% Fe2O3. On the basis of the samples submitted, silica from 5 of the 6 deposits studied has potential as source material for glass sand. Foundry tests showed that 3 of the 4 samples submitted for evaluation have potential as foundry sand."

Jan 1, 1987

By David R. Ray, Arthur J. Pyron, William F. Thomann

"Proterozoic plutonic and volcanic rocks in the Franklin Mountains, west Texas, exhibit similar values in average concentrations of equivalent uranium (eU), equivalent thorium (eTh) and equivalent potassium (eK) when compared to Proterozoic igneous rocks of the southwestern United States. Variations of radioelement concentrations for all Proterozoic igneous rocks of the Franklin Mountains range from 1.1 to 357.1 ppm U, 2.1to 259.1 ppm Th and 2.5 to 10.2% K.Recent geologic mapping of the Proterozoic rocks in the Franklin Mountains revealed several types of granites previously unrecognized, and these warrant further exploration for U and Th resources. Fluorite-bearing granites and riebeckite pegmatites are likely targets for further investigation, as these have the highest U and Th concentrations compared to all other granites.IntroductionThis is a summary of the paper presented at the CIM Geology Division Uranium Symposium held in Saskatoon, Saskatchewan, September 8-13, 1981. The purpose of this study was to evaluate the distribution of uranium (U), thorium (Th) and potassium (K) in Proterozoic igneous rocks in the Franklin Mountains, and compare radioelement concentrations to Proterozoic igneous rocks elsewhere in the southwestern United States.General GeologyThe Franklin Mountains are located in west Texas and southcentral New Mexico, within the southeastern part of the Basin and Range Province o f the southwestern United States. The Franklins are a complex fault-block range trending northsouth with strata dipping toward the west. Ages of the rocks range from Proterozoic to Cretaceous. Harbour (1972, 1960) provides the most recent account of the general geology of the Franklin Mountains. Most recent accounts of the detailed Pro Proterozoic geology o f the Franklins are provided by Hoffer (1976), Thomann (1980, 1981), Pyron (1980), Ray (1982) and McCutcheon (1982)."

Jan 1, 1983