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By R. W. Lawrence, J. B. Austin, P. B. Marchant, J. T. Chapman

"Biohydrometallurgy is the use of microbiological reactions to assist in the extraction of metals. Bioleaching offers an innovative alternative for the treatment of refractory gold-bearing sulphides (1). One particular cost advantage of bioleaching is the potential to enhance gold extraction after only partial sulphide oxidation. The apparent selective sulphide oxidation is more prevalent where the refractory gold is associated with arsenic and antimony (2). Oxidation of the pyrite minerals and arsenopyrite at different oxidation rates can result in high gold extractions at relatively low levels of sulphide oxidation due to preferential, or selective, oxidation of the arsenic mineral. Arsenic extraction rates in a bioleach process typically exceed the iron extraction rate. Therefore, in concentrates in which the precious metals are principally associated with arsenic minerals, liberation of the precious metals may be possible with high recoveries and less than complete sulphide oxidation. The capital and operating cost advantages of partial sulphide oxidation are inherently obvious and are briefly illustrated herein.Ultra fine milling (UFM) is another alternative for enhancing gold extraction from refractory sulphides. Some refractory gold is simply particle size sensitive and UFM can improve gold extraction without the need for costly sulphide oxidation (3). Fine grinding technology for enhanced gold recovery developed by Metprotech, South Africa, is currently producing 2 - 4 micron ground product using a single pass, stirred ball mill. Proper application engineering has ensured that plug flow occurs through the mill. Plug flow from the feed inlet at the bottom of the reactor ensures that short circuiting of feed does not occur and as a result, the mill product does not require classification.A combination of UFM as a pretreatment for bioleaching could provide a mechanism for enhanced selective sulphide biooxidation and/or increased overall gold extraction following biooxidation due to improved particle liberation ( 4). The effect of UFM on biooxidation costs will be discussed. Only sulphide concentrates are considered herein for application of UFM in combination with biooxidation."

Jan 1, 1990

By S. Yalcin, S. Chang, M. Oliazadeh

"Present day ferrochrome demand is on the rise, as more steel products require this material as an alloying element. It is projected that the stainless steel demands will grow by about 6% per year to 2020; as such, the supply and demand for ferrochrome appears promising. Chromite is the only economical source of chromium, which is commonly upgraded by gravity techniques and used to produce ferrochrome. The gravity separation methods are not able to treat particles finer than 0.1 mm efficiently. Therefore, in processing chromite ore, a large quantity of fine tailings; some 25% of the mineral values have been reported as lost in slimes during processing. The chromite content of these tailings is typically 10-20% Cr2O3 and accumulates into a large stockpile/pond. A chromite concentrator project in Canada, considers a cut-off grade of 20% Cr2O3. This indicates further processing might be required to recover chromite from low-grade stockpile when the high-grade deposits are depleted. Low-grade ores normally indicates disseminated chromite ores with low liberation and consequently fine comminution. This paper reviews the techniques for treating chromite from fine material and provides some examples of successful operations in fine chromite recovery. Beneficiation techniques for fine particles include gravity separation, magnetic separation, and flotation. Several case studies are presented here that used one of the aforementioned methods or a combination of them.INTRODUCTIONChromium is one of the essential element and important material for production of ferrochrome alloys and special steel. Chromium occurs as chromium spinel, a complex mineral with chemical composition of (Fe, Mg) O (Cr, Al, Fe)2 . Chromite ore is exclusively found in ultramafic igneous rocks. Major associated gangue minerals are talc, quartz, serpentine and iron oxides.Chromite is a heavy mineral, with a specific gravity in the range of 4-5, relative to associated gangue minerals in a chromite ore, which makes gravity concentration a suitable process. This is generally true for coarse particles, ranging from 1 mm to 100 mm. Typically, the ore is prepared by crushing (and grinding if necessary) to a suitable size and the material is sized for the appropriate beneficiation techniques. It is common to use heavy medium separation (e.g. dense media drums, dense media cyclones, or Dyna Whirlpool) to treat coarse size and intermediate size particles (1 – 100 mm). Intermediate and fine particles (20 – 0.1 mm) could be treated by jigs. Spirals, shaking table, sluices, and Reichert cones can upgrade fine particles. Figure 1 shows a typical chromite beneficiation process flowsheet. As can be seen, it consists of three major sections: feed preparation (crushing, screening), de-sliming, and concentration. Beneficiation recoveries range from 80 to 85%, although the ratio of upgrading is only 2 to 3."

Jan 1, 2013

By J Petrolito

"To improve the economic and environmental performance of underground mining operations, it is common practice to return the metallurgical waste material underground as backfill. This is called hydraulic backfill and is used to fill the mined-out areas not exposed to further mining. To provide mechanical support for underground mining a cemented backfill is normally used. Since its introduction, cemented backfill has allowed mining companies to increase their ore extraction and improve working and environmental conditions. However, cemented backfill represents up to 50% to 85% of the mine’s operating cost compared with 5% to 10% for hydraulic backfill.These economic pressures have led mining companies to carry out research into the partial or total replacement of cement using a variety of alternative materials. The materials are generally site-specific materials, readily available, and cost-effective relative to cement. Such investigations have led to reduced cost of many backfill mixes while maintaining, and in some cases improving, the physical properties of cemented backfill. However, in most cases, cement still remains an integral part of any binder combination. This can be attributed to the alternative binders investigated being either non-binding materials or capable of generating only low-strength backfills, except when used in conjunction with cement.This paper provides a historical review of backfilling binder materials and their characteristics. In particular, it reviews the many alternative materials that have been proposed to either partially or fully replace Portland cement as the main binder."

Jan 1, 2005

By G. Mihaylov

The basic principles of systems analysis are presented. A model has been designed to determine the location of the Main opening in the presence of three or more stock orebodies. The necessary condition is to find a location for the Collection point which is an essential element in opening up each orebody (stock). The model takes into account the total transportation costs from each orebody to the collection point. Three variants are considered provided that the orefield is divided into two levels at least: - a collection point is designed at every level; - a collection point is designed at the lowest level, i.e. where the so-called concentration level is formed; - the collection point is located at a point between the planes of the two levels; The effect of the depth of mining and mine output on determining the method of opening up the orefield has been analyzed by considering a vertical and/or an inclined shaft. A function has been constructed which allows for investigating in both a static and dynamic regime the capital costs for constructing the whole complex including the shaft. The character of the function does not allow finding the global limits. Therefore, it is of particular importance to specify the limiting conditions. In this case these are the flow capacity of development openings and the shaft. The basic groups of cost and natural indicators for assessing and selecting an optimal technological solution have been systematized. The combined analysis of the two approaches to finding the location of the Main opening and determining the mining method enables us to determine when an inclined shaft is best, and the depth at which opening by a vertical shaft has no alternative. Thus, a transition area is formed for which the selection of an optimal technological solution requires an analysis similar to that proposed in this paper. Conclusions are finally drawn which serve as a basis for determining the areas of application of each opening, including the use of a combined scheme of mine opening. The approach presented can be used in both investigating the possibilities for opening up new orefields and reconstructing existing ones in cases when the mining operations go deeper.

Aug 1, 2013

By Yannis Faitakis, Simon Bergeron

"This paper describes the implementation of model-based predictive control (MPC) on a vacuum heat exchange network at Alcan's Jonquière Complex. The plant produces alumina from bauxite ore using the Bayer Process.The MPC system was commissioned in 2000, by Alcan personnel and an Advanced Control Engineer from Matrikon. The application has achieved an uptime of about 97%.The benefits were:• Energy savings : 0.5% - live steam reduction• Reduced variation: standard deviation of key temperatures reduced by 50%• Plant throughput increase: via the automated use of surge tanks• Decreased supervision time: by plant operations• Faster and easier diagnosis of problems and plant constraintsThe benefits were higher than expected with a payback time around 12 months.The purpose of the vacuum heat exchange network is to recover energy from a hot stream (pregnant liquor) by transferring it to a cold stream (spent liquor). The system is very tightly coupled, and the main constraints of the multivariable control system are: • to control the many tank levels involved in the process (8 tanks)• to maximize the temperature of the heated streams (spent liquor)• to prevent cooled streams (pregnant liquor) from cooling to below a minimum temperatureThe system includes 9 manipulated, 14 controlled and 2 feedforward variables.The advanced control implementation runs under an expert system environment and provides added flexibility to the overall control strategy."

Jan 1, 2003

By E. Lampard

Mines must manage variability from within their operations as well as those imposed upon them by external forces. This paper provides a framework to manage both internal and external variability through the review and adjustment of the mine plan where applicable to minimize lost production and maintain operational safety. KEYWORDS Mine plan, mine schedule, variability, safety, lost production, float days, update, adjustment, operations, Maintenance

May 1, 2011

By Andre L. De Vegt, Cees J. N. Buisman

"For the past eight years Paques has been engaged in the development and installation of treatment systems based on biotechnological processes to remove sulfur compounds from water, air and gaseous streams. Metal and sulfate can be removed from mine waters using two biological steps:1. Sulfate reducing bacteria convert sulfate to hydrogen sulfide (H2S). The H2S reacts with the dissolved metals to form insoluble metal sulfide precipitates.2. Sulfide oxidizing bacteria convert excess H2S to elemental sulfur.Paques has designed and installed a groundwater treatment system for the Budelco zinc refinery in the Netherlands to remove metals and sulfate as described above. The system has been operating since May 1992 treating a flow of 5000 m3/d. A test program, utilizing a large scale pilot plant, will commence in November 1995 at a copper mine in the U.S.A. to remediate groundwater contaminated with sulfate and heavy metals. Biological reactor performance and selective metal recovery will be studied. Sulfur dioxide (S02) can be removed from fluegas using a combination of scrubber technology and the two biological steps. The end product is elemental sulfur. A commercial size biological desulfurization plant has been in operation since July 1994 treating 50,000 Nm3/h of fluegas. Sulfide can be removed from mining waters or anaerobic wetlands using a THIOPAQ® Sulfide Oxidizing reactor.lntroductionSulfur can be present in various water and gaseous streams in a mining operation. In many cases the sulfur compounds have to be removed because of toxicity, corrosive properties and contribution to acid rain. Mine waters can contain dissolved heavy metals. An efficient way to remove dissolved metals removal is the precipitation as metal sulfides. Paques has developed and installed treatment systems based on biotechnological processes to remove sulfur compounds from water, air and gaseous streams. These treatment systems biologically convert sulfate to hydrogen sulfide (H2S) which can be used in heavy metal precipitation. Excess H2S is biologically converted to elemental sulfur. This paper provides an overview of these newly developed biotechnological treatment systems."

Jan 1, 1997

By Norman O. Lotter, Simon Yu, Tony Deng, Antonio Di Feo

"Xstrata Nickel’s Raglan operation, northern Québec, achieved paymetal recovery gains in 2007 by switching the collector from potassium amyl xanthate to sodium isobutyl xanthate. To follow on from that position, Xstrata Process Support (XPS) and A. Di Feo of Raglan proposed that a mixed xanthate system be formulated and tested towards further performance gains. This work selected ethyl and isobutyl xanthates as the candidates. An optimized ratio of these was to be determined by a designed experiment. This suggestion was based on an abundance of published literature that reported mixed collector systems outperformed single ones. A partial factorial design was used to configure four test conditions and a mid-square for the flotation work. The testwork was conducted using High Confidence Flotation Testing. The raw data were analyzed and modelled by Design-Expert software. An optimum ratio of 1.6:1.0 by mass (1.32:1.0 in molar terms) of isobutyl:ethyl xanthate was determined. Without this modelling, the optimum ratio would have been missed. Results obtained with the new baseline flotation test using this optimized mixture showed a 3.3% Ni recovery gain in the bypass concentrate and more than 1% recovery gain for PGM over the single PIBX system. This system has been proposed for and accepted as a plant scale trial in the first half of 2010.INTRODUCTIONConcentrator OperationsMilling and flotation of sulphide ores ahead of smelting and refining is a common practice. This choice results in the larger and more variable paymetal loss in the production of saleable metal [Cramer, 2001]. For this reason, it is common to see an ongoing performance improvement programme at the operations level. When successfully performed, these e"

Jan 1, 2010

By K. G. Stowe

"1. Introduction Mattagami Lake Mines and Orchan Mines operate zinc copper mines adjacent to one another near Matagami, Quebec, situated approximately 300 kilometres north-east of Rouyn-Noranda. (Note: Mattagami Lake and Orchan are in the process of being absorbed by Noranda Mines and will become Noranda Mines-Matagami Division). Both concentrators began operations in late 1963.The Mattagami Lake concentrator is capable of treating up to 3,500 metric tons per day of massive sulphide ore with a mineral composition shown in TABLE 1. The Orchan concentrator consists of two separate circuits - the Orchan circuit with a capacity of 1,050 metric tons per day and a custom circuit with a capacity of 825 metric tons per day. Throughout its operation, Orchan has custom milled ore from several smaller deposits in the Matagami district."

Jan 1, 1979

By H. R. Bullis

"Abstract- This paper discusses what appears to be a flaw in the current belief, widely held within the gold mining industry and investment community, that the creation of ever-larger gold mining companies is desirable for mining companies and investors alike. The basis of the flaw is in the underlying assumption that very large gold producers can discover or acquire new gold deposits of the size necessary to replace extracted reserves on a year-on-year basis over the intermediate to long term (i.e., five to ten years and longer). As it will be noted in this paper, deposits of the size required to replace annual production of very large gold producers (VLGPs) are relatively few in number. Although discoveries of new gold deposits continue to be made, they are predominantly in the 0.5 to 2.0 million ounce range. Data represented here suggest that, although gold production from individual VLGPs continues to increase, the change is due to merging or acquiring other companies or projects with similar reserve life profiles. Therefore, the reserve life profiles of VLGPs have remained flat or have decreased.The paper will note recent changes in the industry and the remarkable increase in gold production reported by individual mining companies over the past several years. It will provide an overview of the geographical and geological distribution of known gold deposits and the range of deposit sizes. Finally, the paper will discuss the challenges that face large gold producers in replenishing their reserves that are being depleted at an ever-increasing rate. Unless otherwise stated, all production and reserve data have been taken from company documents in the public domain or publicly available information."

Jan 1, 2001

By V. K. Gupta

"Effect of mill speed on grinding kinetics has been investigated with a view to using mill speed as a manipulated variable for control and optimization of the grinding circuit performance under variable ore hardness conditions. Dry and wet batch grinding tests were carried out for quartz, limestone and cement clinker in a 29 cm x 18 cm laboratory ball mill. Mill speed values selected were 55 and 70% critical. In the case of dry grinding operation, with an exception of quartz, for all materials studied the effect of mill speed on grinding rate was found to be insignificant for sizes finer than about 35 mesh. In the case of wet operation, grinding rates were found to be approximately proportional to the mill speed. The implications of these results are discussed.INTRODUCTIONIt is well known that mill speed has a considerable effect on the grinding rate of the ore particles l-3 For particles of a given size there is an optimum speed at which mill grinding efficiency is maximum (i.e. specific energy consumption is minimum) 3•4. Mill speed can, therefore, be effectively used as a manipulated variable to deal with the disturbances arising from changes in the size and hardness (grinding rate) of the ore being fed to the grinding circuit. Though, it has not been the case so far, with recent developments of ring motors2 and frequency inverters5 a continuous variation of mill speed has now become feasible, and it is expected that mill speed control will soon be a regular feature of plant operation2,6."

Jan 1, 1985

By G. Deschênes

The Eleonore property (James Bay district of Northern Québec), owned by Goldcorp, hosts gold within stockworks of quartz-tourmaline-arsenopyrite veins and veinlets contained within microcline (potassic alteration) and brown tourmaline replacement zones. The preliminary investigation to recover gold from drill core samples showed that flotation followed ultrafine grinding and cyanidation is the preferred option. Ultrafine grinding is used to liberate gold finely disseminated in sulphides. The flotation concentrate used in the investigation had a P80 of 11 microns and contained 75.5 g/t Au, 5.0 g/t Ag, 10.3% S, 0.07% Cu and traces of antimony (0.03%). The mineralogical characterization showed the presence of 65% gangue minerals, 23% pyrrhotite (hexagonal), 9.6% arsenopyrite, 2.2% pyrite and 0.08% chalcopyrite with trace amounts of galena (0.05%) and sphalerite (0.03%). Gold was in the form of native gold and electrum. A straight cyanidation conducted at pH 11.0, DO of 3-5 ppm, 2000 ppm NaCN, 20 C 72 hours extracted 97.0% Au and consumed 31.5 kg/t NaCN. The equilibrium was reached within 24 hours. Increasing the dissolved oxygen to 16 ppm reduced the consumption of cyanide to 22.0 kg/t with 96.7% Au extraction. A 16-h pretreatment with 2.0 kg/t lead nitrate, followed by a 24-h cyanidation reduced the cyanide consumption to 6.0 kg/t with a gold extraction of 97.5%. Increasing the lead nitrate addition to 6.0 kg/t reduced the cyanide consumption to 1.2 kg/t by passivating the surface of pyrrhotite. The gold extraction remained at 97.0%. It was possible to decrease the concentration of cyanide to 800 ppm NaCN without compromising gold extraction. The associated cyanide consumption was 1.2 kg/t NaCN. The passivation of gold grains was an issue when the cyanide concentration was decreased to 550 ppm NaCN. The gold extraction decreased to 92.2%. The average lime addition for all the tests was 7.4 kg/t and the average silver extraction was 80%. Efficient cyanidation of an ultrafine sulphide concentrate with only 800 ppm NaCN represents an advance in leaching of ultrafine sulphide concentrates.

Jan 1, 2013

By G. Major-Marothy

"INTRODUCTIONA beneficiation plant to upgrade earthy iron ores is being installed by the Iron Ore Company of Canada in Sept-Iles, Quebec.The crude ore for the plant will come from the Knob Lake area, located in Northern Quebec, where mining has been in progress since 1954.Some 170 million tons of ""direct 11 ores have been mined and shipped in the past eighteen years, with an overall silica content of 7 to 8 percent. No beneficiation - except the reduction of the top size to about 2 inches - took place during this period, but extensive research work has been carried out, in order to find a feasible process.Since “washing"", gravity separation, roasting and magnetic separation experiments did not indicate positive enough results, flotation was tried and found to be feasible.The beneficiation complex being installed will consist of grinding, flotation, dewatering - i.e.: thickening, filtering and drying - balling feed preparation, and agglomeration - i.e.: balling and induration.This process will result in a product superior to the ""direct shipping"" ores and will also expand the ore reserves by allowing material with a higher silica content to be utilized.THE OREThe Knob Lake ores are grouped by mineralogical composition and by grade.BLUE ORES, representing roughly two thirds of the reserves, carry their iron content mostly in the form of massive, blue hematite and martite.YELLOW ORES, about one sixth of the reserves, have predominantly hydrated iron oxides as the iron bearing minerals.RED ORES, the other sixth of the reserves, carry the iron mostly in the form of red hematite.The gangue in all ores is mainly quartz, but aluminum oxides and hydroxides are also present, particularly in the red ores."

Jan 1, 1972

By Chuck Edwards

"The development of two large, high-grade uranium projects has been proposed for northern Saskatchewan. The projects are now under review by a joint federal/provincial panel. The McArthur River project is owned by the McArthur River joint venture (Cameco Corporation 53.991 %, Uranerz Exploration and Mining Limited 29.775%, Cogema Resources Inc. 16.234%) and operated by Cameco. Cameco proposes to mill the McArthur River ore at the existing Key Lake mill. The Cigar Lake project is owned by the Cigar Lake joint venture (Cameco Corporation 48.75%, Cogema Resources Inc. 36.375%, Idemitsu Uranium Exploration Canada Ltd. 12.875%, Korea Electric Power Corporation 2.000%). The Cigar Lake joint venture proposes that Cameco operate the Cigar Lake mine for the owners, and that Cogema mill the ore at their McClean Lake mill, now under construction. Proposed systems provide safe mining and ore handling, and allow the milling of McArthur River ore at Key Lake and Cigar Lake ore at McClean Lake. This avoids the disturbance of new sites for milling and tailings disposal.The McArthur River and Cigar Lake deposits are the largest known high-grade uranium deposits in the world. The ore grades at McArthur River and Cigar Lake are at least 100 times those that are mined elsewhere. It is proposed to mine these deposits using remote underground mining methods, to crush and grind the ore underground, to pump a thickened ore slurry to surface, and to transport it in special containers by truck to the mills.During all mining and processing stages the ore is fully contained. Radon emission is controlled by positive ventilation of all process vessels, with the process ventilation backed up by the primary mine ventilation system. Radiation is controlled by heavy steel pipe and vessel walls, concrete walls and sometimes lead sheeting."

Jan 1, 1997

By H. A. Schnell

"The Phase I Mill at Cluff Lake in northern Saskatchewan was designed to accommodate a very specific high grade uranium ore. As this high grade ore was depleted, construction was undertaken on a new mill circuit (Phase 2) to process a more conventional lower grade ore deposit.The Phase 2 Mill expanded the ore throughput to 800 tonnes per day. The new process employs crushing, grinding, leaching, counter—current decantation, solvent extraction and magnesia yellowcake precipitation and drying to produce a yellowcake concentrate.This paper discusses the process decisions required for this expansion and outlines the new mill circuit. Included is a review of the start up in August, 1984.1.,INTRODUCTION The Cluff Lake Uranium deposit is located 720 kilometers north of Saskatoon in northern Saskatchewan. A high grade deposit was discovered in 1970 with the first production taking place in 1980 after lengthy hearings. The original mill (or Phase 1 Mill) was specifically designed to accommodate this very high grade ore for a very limited plant life.Pilot plant work was carried out in 1981 on the other lower grade ores found in the immediate Cluff Lake area. Feasibility and process selection studies were then undertaken as the Phase 1 mill was nearing its end of life. These studies were completed in March 1983 when approval was given to construct the new Phase 2 mill."

Jan 1, 1985

By Paul Hughes

This paper summarizes current underhand paste backfill mining as applied to Goldcorp Inc?s Red Lake Mine, Newcrest Mining Ltd?s Kencana Mine and the Lanfranchi Nickel Mine as well as ongoing research at the University of British Columbia?s Norman B. Keevil Institute of Mining Engineering. The Red Lake Mine utilises a 6 m wide undercut with a 2 MPa unconfined strength paste in order to selectively mine its high grade, high-stress orebody. Kencana?s orebody is located in a poor rock mass environment. An 8 m wide, 2 MPa unconfined strength paste sill allows a safe working condition for its high grade orebody. The Lanfranchi Nickel Mine utilises a cable-supported, 2 MPa paste, with 12 m span sills, to optimize the mining of the squeezing rock mass. This paper details the rock mechanic hazards, ground support techniques, paste practices and unique conditions for each mine. Further, an update of ongoing research at the University of British Columbia, focussing on the optimization of safe mining in underhand cut-and-fill mining in high-stress/weak rock masses, is presented.

May 1, 2011

By C. Patel, George H. Hope

"INTRODUCTIONThe Hemlo area is located in Northern Ontario, some 350 km east of Thunder Bay along highway 17. Hemlo is Canada's largest, recent discovery of gold ore. Three companies, Lac, Noranda, and Teck are currently developing mines and processing facilities in the area. A combined milling capacity of approximately 10,000 tonnes per day will ultimately be available in the area. The gold yield will be between 750,000 to 1,000,000 ozs. per year. This will slightly more than double Ontario's gold production but only represents 2.5% of the total world gold production.Lac Minerals Ltd. is substantially committed to the gold mining business. Lac currently operates four (4) gold mines and mills and produces 300,000 ozs. per year. Lac's Hemlo development will put the company in the position of number one gold producer in Canada.The Lac Hemlo ore body has an estimated reserve of 37 million tonnes at 6.2 g/t. The mining will take place at an initial rate of 3,000 t/d commencing in the first quarter of 1986. The mine and plant will be expanded to 6,000 t/d in 1988. At this capacity, Lac will be the largest producer in the Hemlo area.1. GEOLOGYThe orebody is divided into two (2) zones for mining purposes. The 'A' Zone is to the east and dips to the Northeast from surface to 300 m below surface. This zone has a strike length of 250 m to 300 m at an average width of 20 m. The upperpart of the zone will be mined by open pit and the remainder from underground. The 'B' Zone, to the north west of the 'A' Zone, although lined up is not necessarily an extension of 'A' Zone. 'B' Zone has a known strike length of 600 m and is open to depth and along the strike. It is over 70 m wide in the core. This zone holds most of the property's known reserves. 'C' Zone with narrow gold zones is under exploration in the south west portion of the property."

Jan 1, 1985

By H. Wyslouzil

"The Kelsey Centrifugal Jig is a gravity separation device which utilizes the principles of the conventional mineral jigs but with the additional feature of being able to vary the apparent gravitational field. This is accomplished by utilizing the specific induced movements of the conventional jig within a centrifuge.Pilot plant trials have been underway since July 1988 to evaluate the potential applicability of this technology to various process streams at Rio Kemptville Tin.The grade/recovery relationships for samples treated in the jig were consistently better than results obtained from tables. Recoveries of more than 90% were attained at concentrating factors ranging up to 20. When compared with electrostatic separators, the jig achieved equivalent metallurgical results in a two stage process, without the expensive requirement of filtration and drying. Recoveries exceeding 96% were achieved with final concentrate grades averaging higher than 60% tin.INTRODUCTIONMineral jigs have been used for hundreds of years. Many different types have been developed, and each one has numerous variations. In recent years, the idea of applying centrifugal forces to conventional jigs has been put to practical tests and today a number of units are in commercial production. In principle, by enhancing the apparent gravitational field, it should be possible to recover finer particles by overcoming viscosity and surface effects which tend to limit existing gravity devices. In addition, it may be possible to use gravity to separate minerals with much smaller specific gravity differences."

Jan 1, 1990

By R. Venugopal, Saxena. V. K., U. Sankrityayan

"There is a dire need of a technology which can break the mineral interfaces to enhance liberation with minimum energy consumption in mineral beneficiation processes. Understanding of mineral crystal system and interfaces is essential to develop such methodology. So, this study is carried out to find out the role of mineral crystallography in milling and liberation. Data related of mineral crystallography, Vickers Hardness Number, Bond Work Index and beneficiation processes were collected from literature and detailed statistical analysis has been carried out to establish a correlation among these parameters. It was found that atomic weight and cell parameters shown statistically significant relationship with Vickers Hardness Number of minerals, which controls the milling characteristics of ores. Liberation characteristics of five different kinds of ores were analyzed considering the types of interfaces between valuable and gangue minerals. It was found that phase interfaces play vita role in liberation and coarser liberation can be achieved in the minerals of incoherent crystal systems like manganese and zinc ores. Study of this subject can help us to understand the effect of stress wave propagation in solids and its effect on crystal interfaces. It can also help us to design a technology of future to break the mineral interfaces.INTRODUCTIONCommunition is energy intensive mineral processing operation and consumes up to 2% of the total electric power generated in the world. Energy consumption due to comminution is increasing every year as there is growing needs to exploit lower grade ore bodies and process rejects (Blazy et al., 1994; DOE, 2005). Low grade ores require crushing and grinding to liberate the valuable minerals which are further separated from gangues by various beneficiation methods like gravity separation, floatation, leaching etc. Currently, comminution methods are selected based on the reduction ratio and energy consumption but effect of grinding methods on subsequent processing methods specially on liberation also need to be studied to optimize the entire beneficiation process. Most of the attempts made in past to minimize the energy consumption in milling to enhance the mineral liberation were focused on compression mechanism, chemical additives, thermal treatment, microwave treatment and electric disintegration (Kilpelm and Austin, 1982; Norgate and Weller et al, 1994; Bearman et al., 1997; Kumar et al., 2010; Wang el al., 2011; Singh et al., 2012). But the role of mineral crystallography on milling and liberation behavior of minerals not fully explored yet."

Jan 1, 2013

By P. Massi

"Wholly owned by the Coniagas Mining Company, the oldest continuously listed mining company on the Toronto Stock Exchange, St. Andrew Goldfields is located on Hwy. 101 approximately 30 miles east of downtown Timmins, Ontario. Geographically situated on the Porcupine-Destor Fault, the mine and mill in Stock township is positioned in the heart of Northeastern Ontario's goldmining communities.In July 1988, a 600 ft. shaft, hoist, and compressors were already in place when construction started on a 500 T.P.D. capacity C.I.P. mill and crushing plant.By June 15, 1989 the mill was commissioned and attained the design capacity by August of that same year. Since then, changes to the circuit have increased throughput to 900 T.P.D. from two significantly different ore sources.Meeting the Ministry of the Environment's guidelines is an ongoing commitment utilizing the Inca S02 process in line between the tailings pond and secondary polishing pond.INTRODUCTIONExploration on what is now known as St Andrew Goldfields' Stock Township Mine was carried out ln the early l960 1 s by Hollinger Mines Ltd. The property was acquired by Quebec Sturgeon River Mines 1n 1973 and the main shaft was started ill 1979.When St Andrew Goldfields was incorporated 1n 1983, exploration and development commenced to identify the ore body. The company decided 1n 1987 to bring the Stock township property into production after an extensive feasibility study and testwork at Lakefield Research in Toronto. Leslie Engineering was contracted to design, engineer and procure al 1. equipment from the coarse ore bin to the tailings distribution system and backfill plant"

Jan 1, 1992