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By Michel Garon

"Several years ago computer control of the flotation process was successfully implemented at the Matagami concentrator of Noranda Mines Inc. At that time this was the first computer application being installed at this operation. The last few years have seen an extensive use of programmable controllers in areas such as the copper and zinc dewatering circuits and the lime preparation system. With the aid of such instruments, these areas have attained a high level of automation and as a result, a high degree of efficiency. The controllers have proven to be very dependable and easy to use.IntroductionThe Matagarni Division of Noranda Mines Inc. operates two mines in the Matagami District of northwestern Quebec which is located about 650 km northwest of Montreal. The mines ship ores to a central concentrator whose throughput averages 95,000 tonnes of ore per month. Metal contents of the mill feed average 0.9 per cent copper and 5.0 per cent zinc. Concentrate outputs are approximately 3,200 tonnes of copper and 8,200 tonnes of zinc per month.The concentrator commenced operation in 1963. An on-stream X-ray analyzer was installed in 1968, providing analyses of several samples in the flotation circuit without the aid of a computer. In 1974, control of the circuit by computer was added to the system. As a result, zinc flotation efficiency improved considerably and reagent costs decreased through the reduction of copper sulphate consumption.In the past few years it was realized that significant improvements could also be made in other parts of the operation . Copper and zinc dewatering circuits and the lime preparation system were identified as major areas that could drastically benefit from automation. In 1982, programmable controllers were seriously considered as an attractive alternative for achieving the company 's goal. Because of their relative lowcost and their high degree of flexibility, they were selected for implementing automation and, for improving the efficiency in each of these areas."

Jan 1, 1985

By Ronald E. Pike

"PREFACEAt the CIM Annual General Meeting in Quebec City, R.E. Pike of The Toronto-Dominion Bank described how the Canadian banking industry worked with the minerals industry to assemble the large capital requirements for resource projects in Canada and overseas. He developed a profile of the mining industry and Canadian banking, both being important sectors of the Canadian economy and global in context. He also outlined the banking and financing partnership that existed in the past. In looking to the future and challenge that lies ahead f or each industry, Mr. Pike questioned whether the financing practices of the 1970s will be acceptable over the next decade. In that regard, he highlighted the risks that impact on each industry and illustrated how that financing partnership might evolve in response to those forces. The folio wing is a synopsis of his paper.IntroductionThe Canadian mining and metals industry is prominent in the domestic economy and respected internationally. Any measure of its performance illustrates that status-percentage of GNP, employment, exports, capital investment, size relative to other producing nations, etc.Energy, Mines and Resources Canada has forecast that the non-fuel minerals industry will grow in real terms by an average annual rate of 3.0% over the period 1981 to 1991. Using the GNP as reference, that growth in non-fuel mineral production will represent 7.7% of GNP. In translating that growth to capital expenditure, the EMR Canada estimates the aggregate amount at $43.8 billion ('79 dollars) or approximately 12% of total business investment in Canada during the 10-year period.About half of the $43.8 billion will be devoted to the ferrous sector-$8.4 billion in iron ore mining and $12.4 billion in the steel industry. Capital investment in nonferrous s projects is estimated at $13.3 billion, and it is projected at $6.7 billion for industrial minerals. The residual $3.0 .billion is exploration spending."

Jan 1, 1982

By E. Ghali, W. Zhang, C. Su, N. Sorour, G. Houlachi

"Zinc electrowinning process is sensitive to the presence of metallic impurities such as lead ions that could substantially and negatively affect the power consumption and surface morphology. Effect of various concentrations of manganese ions and traces of lead ions up to 2.5 mg L-1 added to the electrolyte during electrowinning have been investigated by electrochemical polarization techniques such as galvanostatic, cyclic voltammetry and potentiodynamic. Scanning electron microscopy (SEM) and X-ray diffraction (XRD) were applied to examine the surface morphology and crystallographic orientation, respectively. Microwave plasma-atomic emission spectrometry has been used to determine the lead content in the deposit and in the electrolyte after electrodeposition. It was found that lead ions added to the zinc electrolyte led to the increase of cathodic overpotential and current efficiencies of zinc deposit. However, addition of Mn2+ into the zinc electrolyte resulted in the decrease of cathodic potential and current efficiency of zinc deposit. It can be stated as example that galvanostatic polarisation at 40?C and 52.5 mA cm-2 in presence of lead up to 0.1-0.2 mg L-1 in the electrolyte and 8-12 g L-1 Mn could be better considered as practical working parameters for electrowinning process.INTRODUCTIONApproximately 85% of the global production of zinc metal is produced by electrolytic process (Sinclair, 2005). The main goal of the zinc plant is to obtain pure zinc deposits with high current efficiencies. Smooth and pure deposits with high current efficiency (CE) are always obtained from pure electrolytes (Alfantazi & Dreisinger, 2001). The presence of metallic impurities in the zinc sulphate electrolyte is a major problem for the zinc electrowinning industry, since decreases in current efficiency, changes in deposit morphology and cathodic polarization are observed for electrolytes containing small concentrations of impurities (Ault & Frazer, 1988; Muresan, Maurin, Oniciu & Gaga, 1996). Low level concentrations as parts per million of these impurities are reported to influence negatively the zinc electrodeposition (Beshore, Flori, Schade & O'Keefe, 1987)."

Jan 1, 2016

By I. Farr, P. Quinn, H. Mani

"Difficulties in the flotation of Inca's ultramafic ores at the Thompson Mill are related to the presence of serpentine slimes and some floatable rock minerals. Serpentine slime coatings depress pentlandite recovery while rock flotation decreases concentrate grade. Carboxy methyl cellulose (CMG) has been shown to promote sulphide flotation and to reject floatable rock.To improve ultramafic ore flotation, CMC addition in the rougher stage was studied in a laboratory scale column. Batch tests were conducted in the lab and followed up with continuous testwork at the Thompson mill. At 0.5-1.0 g CMG per kg ore, pentlandite recovery is increased by 5-10% in rougher flotation, pentlandite/rock selectivity is improved and pentlandite/pyrrhotite selectivity is unaffected.IntroductionAt Inca's Manitoba Division, two distinctly different sulphide ores are processed at the Thompson Mill, primarily by froth flotation. The sulfide ores are contained within the Thompson Nickel Belt and currently there are two operating mines, Thompson and Birchtree. Sulphide mineralization in both mines is simple, consisting of pyrrhotite (both monoclinic and hexagonal), pentlandite, and minor chalcopyrite. Both deposits are hosted by metasedimentary and metavolcanic rocks with variable ultramafic rock associations. Ultramafic rocks are mostly serpentinites: magnesium rich rocks composed predominately of serpentine. There is little serpentinite at the Thompson mine, while mineralization at Birchtree is closely associated with significant ultramafic rocks.Excellent recoveries of pentlandite (Pn) and chalcopyrite (Cp) are achieved at the Thompson Mill when processing ore from the Thompson mine. In treating the ultramafic associated Birchtree ore, nickel recoveries are some 6 to 8% below that achieved with Thompson ore. Furthermore, a typical concentrate from Birchtree ore contains more rock and is higher in MgO content. The presence of rock in the concentrate is related to free floatable rock species as well as rock locked with sulfides. The reason for the difference in flotation responses of these two ores has been the subject of on-going re-search."

Jan 1, 1998

By Laxman M. Amaratunga, Nuri Hrnidi

"Reactive pyrrhotite tailings were agglomerated using the Cold Bond Tailings Agglomeration (CBTA) process with low cost binders; Portland Cement and fly ash. Various chemical additives, were used to retard acid mine drainage (AMD), including acid neutralizers, surfactants, sealants, and bactericides, which were encapsulated into the agglomerates. The choice of these additives was based on their low cost, low toxicity, and environmental friendliness. The main objective of the study was to examine the long term stability of these reactive tailings agglomerates towards AMD when used as aggregates for disposal in underground paste backfill.A carefully controlled kinetic stability test was carried out using specially designed humidity cells suitable for testing pellets. Using the humidity cell test method a 20-30 year AMD test may be accelerated to 50 weeks. During this period the following parameters were monitored: pH, redox potential, conductivity, sulphate, acidity, and dissolved metals.Results indicated that most of the encapsulated additives were successful in retarding AMD, while the control samples produced significant amounts of acid generation. The water soluble surfactant Enviroseal, and the sodium silicate sealant were very effective as a long term retardant for acid generation reactions in the sulphidic tailings when exposed to oxygen, water and iron-oxidizing bacteria.IntroductionThe sulphidic material discharged from concentrators becomes reactive upon exposure to air and/or water. Pyrrhotite (Fe0 S0_1) and pyrite tailings in particular which are discharged from the beneficiation process in the mining and mineral processing activities consist of finely ground particles. Upon exposure to oxygen and moisture, the tailings become reactive due to the instability of the metal sulphide minerals present which undergo spontaneous oxidation, thereby producing a weak sulphuric acid. Therefore, the long-term behaviour of the tailings can have a significant impact on the environment, due to acid mine drainage (AMD). AMD is created during natural atmospheric weathering when leaching action (natural and bacterial) conditions are present in the process wastes discarded during the mining and milling of sulphur bearing ores, and consequently it plays a major role in surface water chemistry."

Jan 1, 1998

By William A. Napier

The Mantaro Phosphate Property is located ~ 250 km east of Lima, Peru in the Andean altiplano near the city of Huancayo, Junin District. The Mantaro deposit is a mineralized zone of phosphatic rock currently defined by surface outcrops, trenches and drilling extending over a strike length of more than 30 km and a width of more than 5 km covering ~27,700 hectares. Mineralization is exposed as three roughly parallel exposures (?mantos?) trending in a northwest-southeast direction along the Mancaspico syncline. Estimated mineral resources1 contained in Manto 1 of the Property (4% P2O5 cut-off) as of February 21, 2010, are: Measured 5.548 million tonnes @ 10.8% P2O5 Indicated 33.975 million tonnes @ 9.9% P2O5 Measured + Indicated 39.523 million tonnes @ 10.0 % P2O5 Inferred² 376.265 million tonnes @ 9.0% P2O5 1Tonnes rounded to the nearest 1,000 and grade to 1 decimal place. 2Inferred material reported with no cut-off and an assumed grade. There is considerable potential to expand the resource base. Mantos 2 and 3 of the deposit have been geologically mapped and exhibit surface mineralization widths similar to Manto 1. Given that the deposit is of marine sedimentary origin, it is reasonable to infer that comparable grades and depth of mineralization are present within Mantos 2 and 3. Based on this assumption, the potential mineral deposits within Mantos 2 and 3 are estimated to be approximately: Manto 2 425-435 million tonnes @ 9% P2O5 Manto 3 280-290 million tonnes @ 9 % P2O5 These potential quantities and grade are conceptual in nature, there has been insufficient exploration to define a mineral resource and it is uncertain if further exploration will result in these targets being delineated as a mineral resource. However, should such conceptual resources be confirmed by trenching and drilling, the Mantaro Phosphate Property would rank as one of the most significant phosphate deposits in the world.

May 1, 2011

By F. Channier

The paper provides the most recent information on the construction of the AP60 phase 1 plant of Jonquiere which is due to be commissioned and started up in the first quarter of 2013. HSE management on site by RTA, and cooperation with local equipment manufacturers and contractors of the Saguenay-Lac Saint Jean region are highlighted in this section. The AP60 phase 1 Jonquiere plant aims to demonstrate of the AP60 technology at an industrial scale and allows the further developments of the AP60 cell. The advantages of the AP60 cell technology in terms of productivity, power consumption and environmental footprint are also briefly presented in the paper.

Jan 1, 2011

By John Lund

INTRODUCTION GOLD-BEARING placers of Alaska and the Yukon Territories are buried under a layer of permanently frozen alluvial overburden, or 'muck', which varies in depth from twenty to one hundred feet. The placers, and the decomposed bedrock layers upon which they are deposited, are also permanently frozen. During the summer, surface muck which is exposed to the sun will thaw, but only to a depth of a few feet. To reach a valuable deposit, the frozen overburden must be either penetrated with shafts, or stripped from the deposit entirely. In either case, before the placer can be mined, it must be thawed. At modern dredging operations, the frozen muck is stripped from the placer by powerful jets of water from hydraulic giants. The placer, and several feet of decomposed bed-rock, are then thawed by forcing water through narrow pipes which have been driven into the deposit. Other methods of thawing frozen placers were used extensively during the early years of the Klondike 'gold rush', but experimentation later proved that creek water at its natural summer temperature was a more effective and economical thawing medium than any other which had been used. EVOLUTION OF COLD WATER THAWING Early prospectors of the far north reached placers by digging shafts through the frozen muck. They then thawed the placer sand and gravel by building wood fires at the bottom of the shafts.

Jan 1, 1951

By J. Lammers, D. Adams

Most pressure oxidation operations introduce oxygen through a large multi-use nozzle in the vapour space of the autoclave. Sparge pipes extend inside the vessel to the point of discharge below the compartment agitator. However, two zinc pressure leach operations have successfully used a bottom entry oxygen sparge system for a number of years. Recently, the concept has been used in a whole ore gold pressure oxidation operation as well as a copper pressure leach plant. Both of the latter operations also introduce cooling water for temperature control through the same sparge pipe. This paper addresses the benefits of the bottom entry system with regards to autoclave performance and the positive effects on agitator performance.

Jan 1, 2004

By Jennifer Abols

In July of 2013, Lundin Mining acquired the 2,000 tpd Eagle nickel and copper project from Rio Tinto. The project had been put on hold earlier in the year and construction was approximately 50% complete. Lundin immediately accelerated project activities and by September of 2014 the project was constructed, commissioned and ramping up. The newly commissioned facilities went on to reach design throughput, recovery and product quality targets by the end of the year. This was accomplished ahead of schedule and below budget despite numerous obstacles including the worst winter on record in the past 50 years for the area. What led to the success of the project, and in particular the commissioning phase of the project? This paper takes a high level look back and provides an overview of the methodology used, key issues encountered and lessons learned to answer this question.

Jan 1, 2015

By Ed McConaghy, P. LaRochelle

The Ambatovy Nickel Project comprises the development of mine, hydrometallurgical processing facilities and necessary infrastructure in Madagascar, positioning Ambatovy as one of the world’s largest producers of nickel and cobalt from lateritic ore. SNC-Lavalin successfully provided comprehensive Engineering, Procurement, Construction Management (EPCM) services for this major capital project which, apart from its sheer size, scope, and engineering design, represented significant location, infrastructure, labour, environmental and climatic challenges. Today, the project has reached 90% of its nameplate capacity and is in its final performance optimization stage. This paper will discuss the significant project management effort that was required to successfully address the complexity of this major project. It will cover the main project management control systems and the engineering tools and processes used during the detail engineering, procurement, and construction phases to achieve optimum project execution.

Jan 1, 2015

By T. R. Stanley

Introduction The purpose of this paper is to interest mining men in the economic possibilities of sand deposits suitable for modern foundry work. Many members of the profession are constantly travelling over the country and it is felt that some knowledge of the necessary properties of foundry sands will enable these men t0 recognize such deposits. It is the experience of the writer that good sands suitable for the varied work of a foundry, such as that of the Consolidated Mining and Smelting Company, at Trail, are almost impossible to obtain from western Canadian sources. It is hoped that the following discussion may arouse some interest in the search for such sands. Sands are used in foundries for two purposes. The main use is for making moulds suitable for receiving hot metal. The mould must retain the? required shape until the metal has solidified. The second use of sand in the foundry is as an abrasive in sand blasting. This is a relatively minor use, and consequently will not be dealt with in detail. Any coarse sand with a high content of quartz will serve this purpose, but round-grained sands are superior to angular types, because they have not the same tendency to shatter on impact with the casting. Sand plays an extremely important part in the successful production of sound castings. This is more evident today than ever before, due to the greater variety of castings and alloys, each of which requires special sand characteristics. Poor sand is responsible for a great many foundry troubles which formerly were attributed to bad? metal; in fact, one investigator goes so far as to say that ninety per cent of bad castings can be traced directly to the sand. Very little of the vast amount of sand distributed over the earth's surface is suitable for moulding. It is the purpose of this paper to describe briefly the properties which are essential in good moulding sand. Sand for this work must be able to withstand the wash and hydrostatic pressure of molten metal, and yet be sufficiently porous to allow the gases generated in the mould to escape.

Jan 1, 1935

By A. A. Bakke

"The Fort Knox gold deposit is hosted within the Fort Knox Pluton, a Cretaceous multiphase granitic body that intrudes the Fairbanks schist. Gold occurs along the margins of stockwork veins and veinlets, quartz-filled shear zones, and along fractures within the granite. The mineralization is generally sulphide-poor and closely associated with trace amounts of bismuth, tellurium and to a lesser extent with molybdenum and tungsten. At present, proven/probable reserve estimates, based on a cutoff grade of 0.39 g/t Au, are 158.3 million tonnes grading 0.83 g/t Au.The Fairbanks schist comprises the bulk of the country rock. The dominant lithologies present include muscovite-quartz schist and micaceous quartzite. Immediately north and south of the main orebody, mixed metamorphic lithologies of the Cleary sequence, a subunit of the Fairbanks schist, have been mapped. Development of local zones of weakness, most likely due to a broad regional warping event, resulted in brittle-ductile deformation and activation of a northwest-southeast shear couple between contrasting lithologies.The Fort Knox Pluton consists of at least three phases based primarily on textural differences. Intrusion of a fine-grained granite was followed by a medium-grained porphyritic granite. The youngest intrusive phase is a coarse-grained, seriate porphyritic granite, the Vogt Stock. The Vogt Stock forms an elongate dome with the axis in the east-west direction. Structures displayed in the overlying units, believed to be the result of doming, include tension fiSSUres and longitudinal fractures.Premineralization fractures, the result of the magmatic doming, provided conduits for initial stockwork quartz vein development and quartz enrichment in shear zones. Weak development of vein-controlled phyl/ic, potassic, albitic, and argillic alteration styles are present. Evidence of a late thermal event resulting in a low-temperature alteration assemblage of calcite + zeolite (stilbite) + chalcedony + clay occupies breccia zones, joint surfaces, and fractures within the granite."

Jan 1, 1995

By J. A. Maritz

"The Hedley and Grant formulation for pillar strength has been widely adopted by the South Africa hard rock mining industry. Today, some forty year later, this formula still provides the foundation for pillar strength calculations when designing for stability and utilising pillars in mine layouts. Various mine specific back analyses were conducted to adjust the equation parameters based on underground observations of failed pillars and comparing to the pillar stresses calculated using numerical models. In all these instances, the influence of shear stresses was ignored. The key to more representative pillar strength estimation is a better understanding of the pillar loading environment.In this paper, special mining cases are evaluated and the existence of shear stresses on the pillar loading environment is illustrated. By revisiting some back analysis data for optimisation of the equation parameters, recommendations were made in respect of the influence of the shear stress component and on how this should be considered in determining the pillar strength.INTRODUCTIONCommon practice on flat dipping, shallow and intermediate depth ore deposits in the South African hard rock and coal mining industry is to exploit the mineral resource using a room-and-pillar layout. The design methodology is mainly based on a Factor of Safety (FoS) approach, where the strength of the pillar is compared to the effective load on the pillar. In these environments, FoS values of 1.6 are applied as it was found to correspond to a probability of failure in coal mines of 0.001. As discussed by Ryder and Jager (2002) page 279: “For lack of anything else more substantial, the figure of 1.6 is generally used in hard rock pillar design in South Africa”. Dependant on the designed purpose of the pillar, the FoS might be selected as high as 2.0 where long term stability is required. The factors directly influencing the FoS include the pillar height (which is linked to the mining height), effective pillar width, a rock-strength factor and extraction ratio."

Jan 1, 2015

By D. Ibarra, J. Jacob, A. G. Guezennec, P. d’Hugues, A. Archane

The lack of adequate gas mass transfer is a rate limiting step in many bacterial leaching processes. Oxygen often becomes a limiting factor because of its low solubility compared to the high demand induced by sulfide oxidation. One way of increasing the solubility of oxygen in water or in media solution is by increasing the driving force, i.e. raising the oxygen partial pressure in the gas stream supplied to the leach pulp. The use of oxygen is a well-known practice in high-temperature bioleaching reactors (above 70°C) whereas air is usually preferred in medium and low-temperature operations in order to avoid too high dissolved oxygen concentrations which could impact negatively the bacterial activity. The purpose of this study was to investigate the use of oxygen-enriched gas in bioleaching reactors at 40°C in order to improve the gas transfer in the system when operating at high solid load (20% and more). Bioleaching experiments were performed on a sulfide-rich tailing waste (pyrite 60%, copper 0.8%, cobalt 0.06 %, gold 1 g/t) using the “BRGM-KCC” bacterial consortia. The reactor used for the tests was designed on the basis of a new bioleaching reactor concept developed by Air Liquide, Milton Roy Mixing and the BRGM which uses floating agitators to inject gases and to mix and to suspend solids in the bioleaching solution. Two types of tests were carried out: with air injection and with oxygen enriched gas mix injection. The aim of this work was to confirm the capacity of bacteria to grow and to dissolve pyrite in this type of bioreactors in oxygenrich atmosphere, and to compare sulfides dissolution rates during both types of bioleaching tests. The results obtained show that the use of oxygen enriched gas mix does not decrease bioleaching performances compared to air injection. High metal extraction yields were achieved (above 80%). No deleterious effect due to oxygen use was observed on the bacteria despite the high level of dissolved oxygen reached in oxygen tests (up to 14 mg.L-1).

Jan 1, 2016

By Andrea R. Gerson, Roger St. C. Smart, Russell Schumann, Joan E. Thomas, Warwick Stewart, Stuart Miller

Sequential Net Acid Generation (NAG) accelerated oxidation testing using peroxide was carried out on single sulfide mineral samples. The calculation of the nett acid production potential (NAPP) assumes that all the sulfide is present as pyrite and that dissolution results in a 2:1 ratio of H+:SO42-. NAPP as kg H2SO4/t is calculated as 30.6 x wt.% S present. To test whether this calculation is valid a NAPP* value was defined in which instead of using the wt.% S present for the calculation the wt.% S dissolved is used. Stoichiometric analysis of the dissolution products of the sulfide minerals as they would occur at pH 7 indicates, that in all cases, the ratio of H+:SO42- of 2:1 is viable. The NAG to pH 7 and NAPP* closely coincided for pyrite, bornite, chalcocite and covellite. For the three Cu-containing minerals, the NAG to pH 4.5 value was 0 as is predicted on the basis of the speciation likely to be present at pH 4.5. Acid generation is due to hydroxylation of the dissolved Cu between pH 4.5 and 7.0. The NAG to pH 7 and NAPP* values for chalcopyrite, arsenopyrite and pyrrhotite were in reasonable agreement but were not for either sphalerite or galena. Zn is not completely hydroxylated at pH 7 as is assumed by the calculation. For galena, surface analysis identified polysulfide formation on of the leach residue polysulfide. The formation of polysulfide is an acid consuming reaction. Therefore under conditions of accelerated oxidation the NAPP value is a reasonable estimate of the possible acid production in all cases except for sphalerite and galena. However, under conditions of less intense oxidation, for instance in columns, waste rock and tailings dumps, the formation of polysulfides is likely to be more widespread thus reducing the correlation between NAPP and the maximum possible acid production.

Jan 1, 2004

By A. M. S. Hussein, B. E. Rogers

Progressively more stringent emission limits and advances in regulations and enforcement have made air compliance more difficult for metallurgical facilities. This has historically required improved collection and treatment of stack sources. Fugitive emission reductions are now increasingly required to achieve compliance. Typically small in magnitude, fugitive emissions have a disproportionate impact on ground level concentrations, can be visible to the public and can influence workplace health and safety. Moreover, these emissions are complex, as they are spread over multiple sources, distance and time and are affected by various activities and conditions. As such it is difficult to find, define and monitor these emissions and their impacts. This difficulty translates to the design of emission reduction measures and operational requirements. This paper presents an overview or ‘heuristic’ of fugitive emissions management for compliance in existing operations. This includes aspects such as the impact of regulations on project scope, approaches to developing a comprehensive emission inventory, considerations for solution development and the plant’s role in sustaining compliance once achieved. Case studies are presented to demonstrate how portions of this approach have been applied.

Jan 1, 2015

By A. McGrath, M. Trovant, M. Haque

Minimizing energy consumption is a priority for many industrial facilities. Metallurgical and gasification processes produce dirty syngas (CO and H2), which offers the potential for energy recovery. However, syngas handling can pose an explosion risk and even minor leaks can result in poor workplace hygiene. Furthermore, many jurisdictions around the world now discourage wet syngas handling due to limited water availability and secondary pollution. Hatch has developed a 100% positive pressure gas handling system that targets these safety hazards without using fresh water. The implementation of a system for calcium carbide smelters at Qinghai Salt Lake Industry Co. (QSLIC) in China is profiled as a case study.

Jan 1, 2015

By M. Zanini, D. Wong

ABSTRACT: Metals market improvements in 2005 drove the desire to increase nickel throughput and the production of nickel oxide sinter at the Vale Ltd. Copper Cliff smelter. Significant difficulties were encountered due to the unreliable quality of Bessemer matte being produced. Concentrate feed changes and the loss of technical and operational expertise during the previous market downturn contributed to the challenge of making the required improvements. Significant progress has been made in the converter aisle by focusing on employee engagement and standardizing best operating practices for a new generation of skimmers. This paper documents these initiatives and seeks to balance the art and science of converting.

Jan 1, 2012

By T. Patching

"THE FOLLOWING notes on coal mining in the Edmonton city area are based, in the main, on scattered reports and accounts as little coal mining has been done in this area for about 20 years and few people are left who were involved in what was at one time an important industry of northern Alberta. Although coal mining is not likely to be permitted in the city area itself in the future, a record of this work is warranted to indicate the ·conditions and problems that may be encountered if underground mining again is proposed in the district in the same geological formations and in similar strata. Coal mining in the Edmonton area developed with the influx of settlers in the latter part of the nineteenth century and the first part of the present century. Probably local coal was used by the Fort's blacksmith as early as 1841, and the early settlers probably were 'gopherholing' into the outcrops along the river banks by the 1870s. The first report of commercial coal mining was in 1880 when a contract was let by the Hudson's Bay Company for the provision of 300 tons. As the population in the city and northern parts of the province grew, many new mines were opened along the river and in the district, and coal was even shipped by barge to points in Saskatchewan(1 ,2) By 1909 the Department of Public Works reported that 'Edmonton stands as the second largest domestic coal producing centre in the province'. It continued to hold this position until 1942. From about 1912 to 1951 annual production from mines in the city and surrounding district ranged between 0.38 and 0.65 million tons, depending on weather, wars and strikes. Production began to decline in 1951 because of the increasing availability and consumption of natural gas which was piped into the city and then provided to other towns and rural areas in the northern part of the province. Taylor noted that between 1880 and March 1970, when the last mine within the city area closed, the total of mines and prospects within the Edmonton city area rose to 153. Most of them were small; not more than 20 produced more than a quarter of a million tonnes during their entire life, and only three produced a million tonnes or more. As described by Taylor ""within the city district workings extend discontinuously from the apex of a narrow triangle near Groat Bridge to the base of the triangle in the Clover Bar area"". The larger mines were in this latter area. Approximate locations of the old mine workings in the city are shown in Taylor's atlas. A number of other coal mines were worked in the Rabbit Hill, Big Island, St. Albert, Namao, Cardiff and Morinville districts around the city."

Jan 1, 1985