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By Y. Shu, A. R. Laplante

"A laboratory Knelson Concentrator (LKC), 7.5 cm in diameter, was used to study the performance of a similar unit at plant scale, 76 cm in diameter, This paper reports on sampling and assaying procedures and problems, and discusses the performance of the plant unit.The sampling procedure was based on the extraction of large samples, 5-15 kg, and their complete processing with a LKC, to characterize size-by-size performance. The procedure yielded reliable estimates of overall gold assays, after blatant assaying errors were corrected. The plant centrifugal unit concentrate was sampled in situ, which yielded very erratic grades, with a large overall negative bias. Due to the very heterogeneous nature of the concentrate, which can be linked to flow patterns in the riffles of the unit, it is recommended that for future testwork, the concentrate be recovered separately, blended and then sampled.The plant centrifugal separator, when operated at a feed rate of 25 t/h, yielded total and free gold recoveries of 41 % and 62%, respectively. This is much lower than previously reported data, and suggest that at 25 t/h, the unit is overloaded. However, recovery did not decrease with time over a loading cycle of 2 hours, which suggests that present loading cycles, generally 1 hour, can be significantly increased. INTRODUCTIONThe study of gold gravity circuits is plagued with the statistical problem of determining the behaviour of particles occurring with a very low frequency. This problem becomes particularly acute with increasing particle size. For example, considering the average weight of a gold flake in the 840-1200 µm (20/28 mesh) to be about 5 mg (Banisi et al, 1991), a reasonably high grade feed of 10 git would require, for a relative error (standard deviation) of 10% on grade, 100 particles, that is 0.5 g of gold, or 50 kg of 840-1200 µm material (a Poisson distribution is assumed). Gy's formula, for the same data, yields a weight of 33 kg, with a shape factor of 0.1 and a size distribution factor of 0.5. Both results are comparable; to obtain the suggested weight of material, a stream containing 5% weight in the 840-1200 µm would require 1 tonne samples (of unsized material). The problem rapidly disappears with decreasing particle size. Still with Banisi et al's data, a gold flake in the 300-420 µm range averages 0.5 mg in weight; the same 10% error on grade corresponds to a weight of 5 kg. Gy's formula for the same size distribution and particle shape factors as above predicts a required mass of 1.4 kg. This would correspond to an unsized sample weight of 10 kg with 14% weight in the 300-420 µm, for example."

Jan 1, 1992

By W. J. D. Stone

"For several years, the existence of a deposit of barite (BaS04) near Brookfield, Nova Scotia, has been known. A very large deposit of strontium (SrS04) near Sydney, Cape Breton, is a more recent discovery. Cape Chemical Syndicate has been formed to work the deposits. A five-million-dollar integrated mineral processing and chemical plant is planned for Sydney to produce various compounds of barium and strontium. The mineral processing plant includes the Jones Wet Magnetic Separator and will produce high-purity barite and celestite. These products will be processed further in the fine grinding plant and the chemical plant. The general project is described, covering the deposits, processing and markets."

Jan 1, 1969

By Jas. E. Thomson

"The Omega mine is in McVittie township in the Larder Lake area, Ontario. Production at the mine commenced in 1936, and, until the end of 1946, ore milled amounted to 1,558,604 tons, from which 207,074 oz. gold and 28,205 oz. silver, valued at $7,585,470, were recovered. The mine workings extend to a depth of 1,975 feet.GEOLOGYThe geology of the mine has been described in detail by C. P. Jenney (l) and the writer (2). The general geology and structure in the vicinity of the mine is discussed on an earlier page in this volume. The rock formations found in the mine workings are indicated on the accompanying plan and section (Figs. 1and2). They consist of an intermingled, south dipping complex of volcanics, sediments, and secondary talc-chlorite schists and carbonates. The carbonatized rocks, which are often referred to as 'dolomite' in the Larder Lake camp, are siliceous-carbonate replacements in the talc-chlorite schist zones. In places they contain green mica, probably fuchsite, and are cut by a stockwork of quartz veinlets and stringers. A stock of syenitic intrusives lies south of the mine workings.STRUCTUREThe main structural features are indicated in Figs. 1 and 2. The formations and orebodies have a pronounced southerly dip, which flattens with depth. The talc-chlorite schist and 'dolomite' zones are part of the Larder Lake fault or 'break' on which there appears to be very large displacement. Evidence throughout the district would indicate that the south side moved up and, in the case of the Omega mine, the thrust may have been distributed throughout the several intensely sheared zones. This slice faulting could account for the repetition of sediments and volcanics. An alternative possibility would be original interbedding of volcanics and sediments."

Jan 1, 1954

By Jean-Christophe Ehrström, Fanny Mas

"A specific test set-up has been developed to study at the lab-scale the high-temperature behavior of aluminium alloys in conditions close to those experienced by thick plates during hot-rolling. A particular attention is paid to reproducing both the stress triaxiality and strain fields that exist in plate during rolling. In addition to standard ductility measurements by means of secondary tensile tests (STT), multi-step plane strain compression (PSC) tests have been performed to mimic hot-rolling pass schedules. Tool and sample geometries were adapted to allow low h/a ratios (h= average specimen thickness, a= tool width) and thus guarantee depressive stress conditions at the center of the specimen during most of the test. Such conditions resulting in porosity opening can then be used to investigate the impact of rolling conditions on damage evolution. The mechanical testing was combined with post-mortem high-resolution ultrasonic scans allowing to detect small amounts of damage and quantify its evolution as a function of processing parameters such as strain and temperature. The test set-up was validated on two 2050-type AlCuLi alloys with different susceptibilities to high temperature embrittlement.INTRODUCTION Aluminium thick plates are widely used in the aerospace industry, mainly as semi-products to be machined for internal structure parts of wings (spars, ribs) and fuselage (frame). The conversion route of such products always involves a hot-rolling step, the conditions of which (temperature, speed, pass schedule) have to be optimized to maximize both mechanical properties and final plate quality. This goal cannot always be efficiently achieved by full-scale industrial trials as (a) cost limits the number of possible test conditions and (b) it is often difficult to change one parameter alone. That is why there is a need for laboratory-scale tests able to mimic the rolling process in terms of temperature, strain rate, stress triaxiality and deformation. Most laboratory hot-rolling mills do not adequately simulate industrial processing routes, especially in terms of heat loss, deformation speed and rolling geometry (maximum roll gap vs. work roll diameter). Concerning thick plate, it is of interest to achieve both similar deformation and L/h ratio (L=arc of contact, h=mean gauge) per pass as the industrial schedule because they are simultaneously involved in damage evolution and void closure during hot forming processes (Saby, 2014)."

Jan 1, 2018

By John C. Leighton, C. O. Brawner, Doug Stewart

"Despite the availability of established, sophisticated methods for planning and designing stable slopes in rock, comparatively little attention is usually paid to the problems of carrying out the excavation. Blasting should be carefully planned to obtain optimum fragmentation as well as steep, stable working and final pit walls. The principal difficulty facing a blast designer is the lack of prior information about the many critical blasting characteristics of the rock mass. The common practise of trial-and-error lasting will eventually lead to a suitable program, but must be repeated time after time in variable geology. This frequently results in many blast-damaged slopes. Recent research indicates that monitoring drill performance will yield a Rock Quality Index which correlates to a Powder Factor value. The drilling can provide a continual supply of data, reflecting changes in the rock mass which enables the selection of an economical powder factor.IntroductionWhen planning to excavate a large-scale rock slope, engineer s will generally spend considerable effort to produce a safe, feasible and economical design. Much time will usually be spent on the field map ping, 'structural geologic studies, surface and subsurface hydrology, determination of strengths and effective friction angles of discontinuities, and finally checking stability with the aid of mathematical and analytical models. If the rock cut is to be along a transportation corridor, the prime justification for this time and effort is to produce a clean, stable slope which will be safe for those travelling or working beneath it. For the walls of an open-pit mine, there is the additional economic incentive to maintain the steepest stable slope angle for the minimum stripping ratio (Stewart and Kennedy)."

Jan 1, 1982

By G. M. Proudfoot

IN June, 1940, a .Longyear 3420 blast-hole drill rig was purchased by the Hudson Bay Mining and Smelting Company for the purpose of experimental blast-hole drilling. The early work showed that the diamond drill could compete successfully with the Leyners then in use and by January 1st, 1941, the diamond drills were placed on contract. They may be considered to have gone into production at that date. Prior to the introduction of the diamond drill, our bench drilling was clone with Ingersoll-Rand S-70 and S-83 Leyner drills. Holes up to 40 feet were drilled, using coupled 1 1/4-inch round steel. Bits with a gauge of 3 inches were used in starting these holes, with a drop of 1/8-inch gauge for each 2-foot run. The holes normally ended with 1 1/2 or 1 5/8-inch gauge, though bits down to 1 1/8 inches were available for holes longer than normal or for holes drilled in abnormally abrasive ground. This method of drilling was very satisfactory and was a distinct improvement on the three-bench method of drilling used previously (1). During 1937 and 1938, when experienced miners were available, Leyner long-hole drilling was carried on with great efficiency. However, in 1939 and 1940, labour turnover in-creased and experienced long-hole drillers became scarcer. This resulted in a marked decline in the efficiency of the drilling operations, with increased costs and less tonnage drilled per shift. EXPERIMENTAL WORK The initial diamond drilling test work was carried on with a Longyear 3420 air-driven blast-hole drill rig, drilling a 1 3/16-inch hole. Tests indicated that diamond drilling could be successfully adapted to local conditions and that overall costs compared favourably with costs obtained with the Leyner long-hole method. Leyner long holes on stope benches usually bottomed at a larger diameter than the 1 3/16 -inch diamond-drill holes. More powder could therefore be introduced into the Leyner holes than could be placed in the diamond-drill holes. To achieve the same fragmentation, additional footage was required when diamond drilling. Diamond-drill holes of larger diameter were indicated.

Jan 1, 1943

By J .. Carignan, Y. Dembele, J .. Guha, A. Fowler, A. Gobeil

The Barlow Lake stratigraphic section comprises the top of the Gilman Formation, the Blondeau Formation and the three sills of the Cummings Complex. Representative samples were analysed to allow characterization of the chemo-stratigraphic profile and geochemical affinity of the six rock facies found in the section. All the mineralized horizons are associated with sedimentary units. Geochemical signatures associated with the mineralizations are either very weak or very localized and shown by increases in K20 around the mineralizations and Fe203 near them. The mineralizations are stratiform, synvolcanic but resedimented and remobilized by regional metamorphism and possibly by hydrothermal activity initiated by the intrusion of the Cummings Complex. The evolution of the sedimentary basin with the decrease of its depth is confirmed by the chemo-stratigraphic profiles. Stratigraphic position in the section is important in order that the chemo-stratigraphy may serve as a guide for mining exploration in the distal facies of the Blondeau Formation.

Jan 1, 1984

By K. Iturralde

Dense transport infrastructure areas lack green facilities, they produce annoying noise and they are big energy consumers. As far as there is no free land or space, adding a vegetal covering to existing infrastructures has become a choice in order to get a more pleasant and comfortable environment. Noise barriers are getting spread all over the highways and sometimes they install photovoltaic panels. Is this just a hype or are there real causes? A green acoustic barrier with an energy generator could be interesting for covering the existing urban transport infrastructures. The specific geometry of the infrastructures requires a personalized product. If we measure the infrastructure environment's characteristics, which are humidity, temperature, sunlight and size, we can create an accurate 3D model. This model would be used for various purposes. For instance, it could be a useful tool to design and fabricate the customized prefab acoustic barriers and plant containers. And it would be helpful for adjusting efficiently the future performance of the vegetation and photovoltaic panels. But the insertion of vegetal elements in infrastructure surroundings can be a tough operation. It requires special skills and techniques. Nowadays, those tasks are performed almost manually and they require a big effort. The prefab containers for vegetal elements would ease the robotic assembly on site.

Aug 1, 2013

By Sunil Koppalkar, Liming Huang

"Since late 1980s the research group led by late Professor Andre Robert Laplante at McGill University undertook fundamental studies and applied research of the Knelson Concentrator (KC) and developed a sustainable method to correctly estimate gravity recoverable gold (GRG). In the early time a concept of ‘free gold’ was used, it was defined as the gold that could be separated and recovered from gold ore by amalgamation. Due to health and safety hazards, its practical use had declined. Around 1994, a methodology using a 3-in Knelson Concentrator to measure gravity recoverable gold was established at McGill University. Through more than 10 years’ improvement and optimization, this technique had become a standard method to evaluate GRG for gold ores, gold plant streams and platinum group metals. Until 2006, about 200 GRG tests had been carried out at McGill University for the ores from all over the world. Recently, a modified version of GRG test (short test) with the advantages of reduced processing time and lower cost had been developed for the mundane needs of gold industry.INTRODUCTIONGold gravity recovery has evolved significantly over the past twenty-five years, mainly because of the advent of the Knelson Concentrator (Knelson 1988, 1990, 1992, Laplante 2000). Now, a series of Knelson Concentrators for different applications are manufactured and used in gold industry all over the world (Knelson Concentrators 2004).In 1980s the Knelson Concentrator gained acceptance from mining industry as an effective device for the recovery of gold from grinding circuit. Knelson Concentrator became the obvious choice because of its design features: a relatively long processing time without concentrate discharge (compared to some other centrifuges, such as Falcon Concentrator and Chinese Yunxi Centrifuge at that time), large throughput capacity, treating a wide size range of material, yielding a high grade concentrate with small mass that could easily be smelted. As more and more Knelson Concentrators were accepted in gold industry, mineral processing engineers faced with challenges: what are the separation mechanisms of the Knelson Concentrator? How much gold can be extracted for an ore if a Knelson Concentrator is installed? For answering these questions, a laboratory scale Knelson Concentrator was created."

Jan 1, 2007

By K. T. Aye

An investigation into the leaching of chalcocite and covellite in cyanide solutions at concentrations typically encountered daring gold cyanidation has been conducted. The results indicate that, for both minerals, the extent of leaching is greater under anoxic conditions, but is still incomplete even in the presence of free cyanide ions, with undissolved copper solids remaining in equilibrium with solution species. The speciation of the sulfur and cyanide reaction products in solution are interpreted in terms of the fast formation of a copper(I)- cyanide complexes in solution and a sulfur-rich copper(I) polysulfide solid phase, which then reacts more slowly to produce thiocyanate ions.

Jan 1, 2011

By M. L. Smith

"In the Phosporia formation of SE Idaho, ore quality varies by depth, seam and distance along strike; phosphorus recovery is sensitive to these variations. Plant ore feed can be blended over time to minimize this fluctuation by using a combination of mine production and inventory stockpile scheduling. Mixed-integer and goal programming has been used to optimize short-term mine production, but the seasonal nature of mine haulage requires the use of stockpiles both at the mill and the mine. The contents of the stockpiles can also be optimized as part of the production scheduling problem, but introducing inventory control introduces a nonlinear term into the formulation. A solution using a piecewise linear formulation is modelled and solved using AMPL/Cplex, and solution strategies are discussed for similar large-scale problems. IntroductionIdaho’s minable deposits of phosphate rock are found in the Meade Peak member of the Phosphoria Formation; two separate zones of this formation are mined with a combined thickness of 30 ft to 60 ft. A generalized stratigraphic section of the Phosphoria section is shown in Figure 1, illustrating the degree of complexity encountered when scheduling mine production. Alteration of the formation results in higher grade ore zones near the surface and variable in P2O5 levels between seams and along strike. This variability is seen as well in other quality characteristics critical to the production of acid and elemental phosphorus. Further, plant recovery and nodule quality are sensitive to variations in ore feed; variability in SiO2, carbonates, moisture, organics (carbon), P2O5 and other constituents of the ore can strongly affect plant recovery. Because ore quality characteristics can vary as a function of the alteration zone (or seam) being mined as well as the depth and distance along the strike of the formation, scheduling of mine production can have a highly significant effect on recovery in the plant. As illustrated in Figure 1, there are nine seams containing 20% to 35% P2O5. Phosphoric acid and elemental phosphorus production requires an ore content of 30% and 25% P2O5, respectively. Variability between numerous beds that are mined and quality variation within a seam due to alteration greatly complicates production scheduling with the result that erratic mill feed characteristics greatly decrease recovery."

Jan 1, 1999

By Lloyd W. Trevoy

Syncrude Canada Limited is presently studying commercial production of heavy minerals from plant tailings of the openpit A lhabasca tar sands operation. When tar sands are processed and upgraded to synthetic crude oil, heavy minerals can be removed f rom the extraction circuit and beneficiated to provide concentrates of zircon and titanium minerals. Calculated recovery rates indicate that crude oil production capabilities of 105,000 barrels per day would provide 26,000 tonnes of zircon and 114,000 tonnes of titanium minerals annually. The titanium minerals concentrate is a mixture of ilmenite, leucoxene and rutile; it could be used as recovered or upgraded to synthetic rutile. With the consrruction of further synthetic crude oil plants in Western Canada, a large reserve of heavy minerals will become available.

Jan 1, 1984

By H. Dean

"Mine hoisting machinery, of which wire rope is an essential part, is developing, in Canada as well as in other parts of the North American continent, into highly sophisticated equipment. Heavier payloads are now required to be raised from steadily increasing depths at higher rates of speed and production. These hoisting conditions have, of necessity, brought about a major change in the production of steel roping wires and the wire ropes manufactured from them.Operators and designers of hoisting equipment are requesting a very much more efficient ratio of strength to weight in hoisting ropes, so that heavier loads may be raised from greater depths by ropes which will provide a predeterminable life span on which the operator may depend.The cost of wire rope to the operator is very low indeed, usually less than $0.01 per ton of payload, but the cost of unexpected shutdown periods, due to rope maintenance or change, can be extremely high -particularly in operations where a continuously high rate of production is an absolute necessity. Mine hoist ropes with a higher strength to weight ratio are now being made and used in some areas of Canadian mining, and a general understanding of the performance we may expect from them is required."

Jan 1, 1968

By Peter A. Cain, John C. Folinsbee

"A 4-ft.-diameter shaft has been drilled to a depth of 2,790 ft. in strong, hard, basic Precambrian rocks at Lynn Lake Manitoba. This was done on the property of Sherritt Gordon Mines Limited for the purpose of ventilating and developing the ""N"" orebody, which is located approximately 4,000 ft. south of the Farley shaft.The drilling was done under contract by two drilling companies. Peter Bawden Drilling (1964) Limited drilled the initial 12:1,4-in. pilot hole to a depth of 3,000 ft. Shaft Drillers Limited of Canada then reamed the hole to 4 ft. in diameter to a depth of 2,790 ft. At this point, the reaming was stopped because of the difficulty in passing a change in direction in the line of the pilot hole. The hole was completed by driving a raise 200 ft. from the 3000 level to the bottom of the 4-ft. hole using an Alimak raise climber.Cuttings from the reaming operation were directed into the 12:1,4-in. pilot hole, where they fell to the 3000 level and were removed by a mucking machine and standard mine development cars.The elapsed time from the start of the pilot hole to the completion of the reaming operation was ten months, with an average advance of 279ft. per month. There was a two-month delay between the completion of the pilot hole and the start of the reaming operation.The 12:1,4-in. hole was drilled in an elapsed time of 45 days from the time of setting up the rig. The average advance per day was 66.8 ft.The reaming operation was done in an elapsed time of 194 days. A one-month delay required to repair damaged equipment left 163 days for the drilling and setting-up operation, for an advance per day of 17.1 ft."

Jan 1, 1966

By Fazel G. Jahromi, Ahmad Ghahreman

"Atmospheric leaching of arsenical sulfide concentrates, and especially enargite leaching, is one of the major challenges within the current hydrometallurgy research area. One of the recent processes developed for sulfide ore treatment is atmospheric leaching in the presence of AF 5, a novel carbon based catalyst, in chloride media. The quantification of different ion concentrations during enargite and chalcopyrite leaching in the presence of AF 5 were measured and the copper recovery rates of 97% and 96% were achieved. The objective of this study is to shed light on the leaching mechanism of the sulfide minerals. Based on the speciation of iron and copper by different titration methods, the AF 5 catalyst first reduced ferric and cupric ions and then oxidized ferrous, cuprous and arsenite ions. The oxidation behavior of the catalyst enhanced the leaching results. The AF 5 catalyst also caused in situ precipitation of scorodite in leaching experiment. The speciation of ions in the chloride media were simulated by Phreeqc software and the dominant species in the leaching solution were determined.INTRODUCTION The hydrometallurgy of copper has been used for several decades and is still under development for more complex copper sulfide minerals due to their refractory nature. Most of the world copper reserves are in the form of chalcopyrite; and enargite is known as a major contaminant. Pyrometallurgical processing of enargite is problematic due to arsenic release and tight environmental regulations (Peacey, Gupta, and Ford n.d.). On the other hand, lower recoveries in the hydrometallurgical processes make most of these processes impractical. Several leaching processes have been developed for the leaching of copper sulfide minerals. Some of the common treatment methods for copper sulfide minerals are pressure oxidation leaching (POX), CESL, Activox, Albion process, atmospheric ferric sulfate leaching, Dynatec, Galvanox, and Hydro copper process (Barr, Defreyne, and Mayhew 2005; Conner and Anderson 2013; D.G. Dixon, Mayne, and Baxter 2008; Hyvärinen and Hämäläinen 2005; Palmer and Johnson 2005; Rohner, Bartsch, and Ngoviky 2012). Generally it is not viable to dissolve copper sulfide minerals at ambient pressure in mild acidic media and dissolution of these refractory minerals requires harsh conditions in the presence of oxidant materials (Biswas and Davenport 2002). Ferric ions have been used as an oxidant in different media in copper sulfide leaching processes. U. S. Bureau of Mines’ Reno Metallurgy Research Center in 1969 (Haver and Wong 1971) investigated ferric chloride leaching of chalcopyrite. Dutrizac and Macdonald (1972) studied ferric sulfate leaching of enargite. Based on this study, 50% of the copper was recovered at 80-85 ºC after 7 days of leaching. Cupric ions can also be used as an oxidant. Parker, Paul, and Power (1981) stated that cupric/cuprous couples are more effective than ferric/ferrous couples in chloride media because they are rapid and reversible. The Hydro Copper process leaches copper sulfide minerals in chloride solution catalyzed by cupric ions. In this process, copper dissolves as cuprous ions in presence of cupric ions (about 10 g/L) (Hyvärinen and Hämäläinen 2005)."

Jan 1, 2017

By T. R. Raponi

"The Williams Mine, Canada's largest gold mine, processes over 6,000 t/d of ore utilizing SAG/ball milling, leaching and CIP, carbon stripping, electrowinning and refining to produce gold dare bullion.Grinding accounts for over 50% of milling costs with grinding media, power and mill liners making up over 75% of grinding costs. Since 1991, several changes in these areas have been made which have significantly reduced costs, while fineness of grind and mill recovery have both been increased.This paper discusses the changes made in the grinding circuit over the past few years and assesses their impact on operating costs and mill operations.IntroductionThe Williams Mine is located in the Hemlo area of northwestern Ontario, about 40 km east of Marathon. It is operated by the Williams Operation Corporation, which is equally owned by Teck Corp. and Homestake Canada Inc. Canada's largest gold mine, the Williams Mine commenced operation in October 1985 at a rate of 3, 000 t/d. Mill throughputs have typically exceeded design, averaging 3, 397 t/d in 1986, and 3,380 t/d in 1987. In August 1988, production was increased to 6,000 t/d. Throughput averaged 4,212 t/d in 1988. Production increased over the next few years averaging 5,477 t/d in 1989, 6,248 t/d in 1990, 5,987 t/d in 1991, and 6,283 t/d in 1992. With a recovery of over 95.5%, production for the first three quarters in 1993 is averaging 6,408 t/d.GeologyThe main ore body is a steeply dipping conformable zone of microcline-quartz-muscovite rock. It occupies the contact zone between hanging wall metasedimentary rocks and foot wall felsic quartz feldspar porpyritic rocks of possible volcanic or subvolcanic intrusive origin. The ore body is enveloped by muscovite schists which are thickest on the footwall side. The schist is enriched with tourmaline, pyrite, and green mica containing vanadium and barium. The ore body contains pyrite and barite which occurs mostly in laminae and bands to veinlets and breccia fillings. The gold occurs as free (native) grains, 20 to 30 µm in size at grain junctions of the microcline and quartz. Visible gold is rare in the main zone. There are commonly two or more discernable layers of economic value which are separated by weakly gold mineralized material which usually consists of muscovite schist, barite rich zones or sulphide poor-microcline-quartz rock and in the upper and western fringes by biotite rich schist with feldspathic fragmental characteristics. Molybdenite is a common constituent as very fine disseminated grains in the microcline-quartz rock and is most prominent with gold enrichment. Stibnite and realgar are two of the more common accessory sulphides."

Jan 1, 1994

By C. Van Den Heuvel, C. B. Van Buuren

BIOX® has been successfully demonstrated over the past 2 decades for the recovery of gold from refractory sulphide concentrates. To date 12 successful BIOX® plants have been installed in 8 countries spanning 4 continents with the 13th installation currently nearing completion in the Philippines. In 2010 BIOMIN introduced the ASTER™ process for improved water balances in commercial BIOX® applications by supporting tightening environmental legislation on water usage and availability in both arid and tropical regions. Previously, a common challenge for BIOX® plants was the inability to recycle thiocyanate containing water due to the low tolerance of the organisms. ASTER™, a biological thiocyanate and cyanide destruction process, renders a non-toxic solution for re-use in BIOX®. Two commercial ASTER™ plants are in operation with the 3rd ASTER™ plant under construction in the Philippines. A 3rd BIOMIN technology, specifically developed for double refractory ores is available for implementation. The HiTeCC process has shown considerable success on ores where competition from disseminated carbon has rendered ineffective gold recoveries. Through HiTeCC, gold is desorbed from the carbonaceous matter by manipulating the ionic strength and temperature of conventionally treated CIL product. One Australian commercial plant utilises HiTeCC while a 2nd installation is set for commissioning 4th Quarter 2014. BIOMIN’s premier solutions for refractory gold processing are described herein.

Jan 1, 2014

By B. Surender Mohan

Mining lignite pollutes the environment and causes ecological damage if unattended. Being conscious of the importance of environmental resources for the well-being of the living system, the Neyveli Lignite Corporation (NLC) adopts the policy of environmental regeneration as the foundation on which productivity is built. NLC started investing on Eco Care and Environment Protection long ago and it continues. Early investment in environmental protection has resulted in steady growth of the Company along with a healthy environment. The future vision of NLC is to be in the vanguard among the contributors to the community, environment and the nation by developing and utilizing industrial and human resources to the optimum. Untiring efforts are being put in mines (NLC as a whole) to maintain and sustain the ecological balance arising due to continuous mining activity. NLC?s Environmental Management System is becoming a symbolic role model for many of the opencast mines in Southeast Asia. NLC has been bestowed with number of prestigious Environment and Sustainability awards in recognition of the mammoth contribution in maintaining a clean and green environment.

Aug 1, 2013

By G. Couture

"The Glencore Raglan mine Concentrator, located in Quebec’s Nunavik region, has been in operation since 1997. The nickel metallurgical plant was originally designed to process 100 tonnes per hour of ore to produce 130 000 tonnes of nickel concentrate per year at 16% Ni. The increased grinding capacity and the discovery of rich ore bodies have increased the annual production to more than 190 000 tonnes of concentrate at 19% Ni. In 2014, the Metallurgical group, with support from BBA, implemented an expert control system on the flotation circuit. This project was initiated to cope with the high variability of the concentrator’s feed. Increasing the plant recovery and final concentrate grade were also important project objectives. The implementation of the expert system ensures better control of concentrate inventory while maintaining maximum plant throughput. This paper describes the successful implementation of the new control system, the control technology selection and the flotation circuit optimization. It will also discuss the performances obtained with this new approach and the improved process stability.IntroductionThe Raglan Mine is located in Northern Quebec along the 62th parallel. It has been in operation since 1997 and was originally designed to produce 130 000 tonnes of concentrate annually at 16% Ni. The increasing of the grinding capacity and the discovery of rich ore bodies have increased the annual production to more than 190 000 tonnes of concentrate between 18 and 20% Ni."

Jan 1, 2015

By H. G. Ames

"The property of Perron Gold Mines, Limited, lies in the southwest corner of Pascalis township, northwestern Quebec. It is approximately 12 miles northeast of the town of Val d'Or.This ground was originally staked by Jack Mathews of Amos in the spring of 1931. Shortly after staking, a quartz vein containing visible gold was uncovered on the south boundary of the property, about 400 feet south of the present shaft. With further prospecting, another spectacular goldbearing vein was found 3,000 feet northwest of the original showing. Subsequently, several programmes of exploration were carried out in search of the main ore zone. By 1933 sufficient ore was outlined to warrant sinking a shaft. The first ore was milled in 1934.Total production from the Perron mine from commencement in 1935 to 1944 has been 317,900 oz. gold having a value of $11,935,106. The average recovery a ton for 1944 was $9.01."

Jan 1, 1954