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By J. M. Janssens, Van Quach Tinh, M. Ruel, R. Leblanc

"Microscopic studies of peat moss indicate a highly porous material. In view of using this natural product as an adsorbing agent for the treatment of polluted water, studies of its adsorbing power for beef extract and alkyl benzene sulphonate solution were determined. The effects of time of contact, particle size, concentration of adsorbent and concentration of pollutant were studied. In the case of beef extract, there is a COD reduction of around 27 per cent; in the case of the alkyl benzene sulphonate, the percentage reduction varied from 72 to 95 per cent. A water treatment plant based on the adsorbing power of peat moss with simultaneous recuperation of the pollutant is proposed."

Jan 1, 1971

By B. E. Halbert, K. T. Himbeault, S. L. Fernandes, R. L. J. Phillips, B. J. Topp

The Beaverlodge uranium mining and milling facilities operated from 1952 to 1981 with about 94% of the ore extracted from the main underground mine and 6% from smaller satellite deposits. Decommissioning work occurred from 1982 to 1985 involving periods of shutdown, salvage and reclamation. Transition phase monitoring, leading to eventual delicencing commenced in July 1985. Over the last 15 years, discharge from the tailings management facility (TMF) and a fresh water stream, impacted during the operational phase by tailings spills, has improved in water quality for most parameters of concern. Loadings to the environment of three key contaminants (radium-226, total dissolved solids, and uranium) have consistently been less than during the operational phase with radium-226 having the greatest variability. Outstanding environmental issues associated with the recovering drainage system formerly used for tailings disposal, are being addressed in an enhanced environmental monitoring program to commence in 2000. Changes in water chemistry and the natural re-introduction of aquatic organisms are issues of concern. In the meantime, application for release from selected satellite areas commenced in 1999. This paper reviews the issues which have arisen during the transition phase, outlines how they have and are being addressed, and provides a comparison of original and current predictions of the recovery process under way at the former Beaverlodge mine site.

Jan 1, 2000

By W. Kunda, J P. Warner

SHERRITT GORDON Mines Limited annually produce 300,-000 pounds of pure cobalt by processing a nickel-cobalt-copper-iron sulphide ore which is mined at Lynn Lake, Manitoba. The ore (containing 0.04 per cent Co) is •concentrated to give a concentrate containing l 1-12 per cent nickel, about 2.0 per cent .copper, and 0.25-0.35 per cent cobalt, which is treated at the Refinery in Fort Saskatchewan, Alberta. The concentrate is leached in an ammoniacal solution of ammonium sulphate using compressed air as the oxidant. After filtration the leach solution contains, besides nickel and cobalt, copper ammines, sulpharnates and salts of various thionic acids. The copper is removed by precipitation as sulphide and the sulphur compounds converted to ammonium sulphate by a combined oxidation and hydrolysis treatment. The purified solution •Containing about 45 gpl of nickel and 1 gpl of cobalt is treated with gaseous hydrogen under pressure at 350°F to precipitate metallic nickel powder, the precipitation 'being terminated when the nickel content of the solution is about 0.8-1 gpl. Only about

Jan 1, 1962

By D. J. King

The development in recent years of a fluidized bed system for use as a counter-current solid ion exchange loading contactor for the purification and concentration of uranium from leach liquors has had a major impact on the design of uranium extraction plants. The main object of this paper is to describe many of the variations that can be applied to the design of uranium extraction plants arising from the use of a fluidized bed loading contactor with a view to providing a guide for the development of a flowsheet for a new uranium extraction plant. The paper has been devoted primarily to the consideration of conventional uranium extraction plants in which a ground ore is leached in mechanically or air agitated tanks. The effect of the use of a fluidized bed loading contactor on each of the plant sections from leach to uranium purification and concentration has been considered. In order to describe most effectively the impact of using a fluidized bed solid ion exchange loading contactor, its use has been compared with the performance that can be achieved with the most common alternative purification processes. The comparisons are made largely on a technical basis, but the effect on the capital and operating costs of the different uranium extraction flowsheets are described qualitatively. In making these comparisons, consideration is given to the effect on the size, number and materials of construction of equipment items and how the quantity and composition of the liquid effluents can be controlled to reduce the environmental problems arising from the disposal of these effluents. The paper also discusses some of the factors that should be taken into consideration in the design of a fluidized bed loading contactor and describes various designs of the elution section that can be used in conjunction with the loading contactor.

Jan 1, 1979

By A. H. Ransom, M. E. Robb

The Salmita Gold Mine is located 240 air kilometres northea t of Yellowknife on the east shore of Matthews Lake. Salmita is one of several gold showings which are regionally associated with the steep easterly dipping contact zone between mafic, intermediate and fe lsic volcanics to the west and overlying finely laminated argillaceous sediments to the east. The volcanic assemblage forms a narrow 2 m to 7 m wide north to northwesterly trend ing Belt of Yellowknife Supergroup Archaen greenstone rocks. The contact zone exhibits exhalite character and the increased sulphide content provides a di stinct VLF-EM , IP/Resisti vi ty and/or magnetic signature. Gold occurs at or near this contact in three main settings: associated with si liceous vein-like structures; with finely disseminated sulphides within ericitic altered upper fe lsic tuffaceous unit ; or with quartz in narrow shear zones within the overlying sediment . The B-Vein is of the former type with free mi ll ing gold occurring within a highly sil iceous zone having an average widtn of 1.69 m. Alteration is intense within the immediately su rrounding host lithologies and includes the development of si lica, carbonate, sulphide, sericite and chlorite with minor graphite, epidote. tourmaline and scheelite. The mineable zone above the 4th level (-200 m) ha an average strike of 160 m and is divided into three south plunging ore shoots separated by the regional Salmita Shear Zone in the south and by a low-grade structural hiatus in the north. Production from the B-Vein commenced in August 1983 based on a diluted in situ ore reserve of 140,000 t grading 26.6 git Au . As of December 31, 1985, a total of 131 ,537 t grading 27.8 git Au have been mined at a ra.te of 160 t per day and a 96.4 percent recovery, to produce 3,522 ,000 g (113,227 ounces) of gold . Current (January I, 1986) reserves total 45,500 t grading 37 g/t Au down to the 10th level (-500 m) .

Jan 1, 1986

By L. C. Lien

Tower crane layout (TCL) is a typical construction site layout (CSL) problem, which is suitable for a wide range of work assignments and site conditions. Tower crane is one of the key facilities for vertical and horizontal transportation of materials, especially for heavy prefabrication units such as steel beams, ready mixed concrete, prefabricated elements and large panel formwork such as machinery and equipment, and a wide variety of other building materials within a construction site. However, it is a difficult combinatorial optimization problem to determine the locations of tower cranes with materials quantity supply and demand for engineers. Swarm intelligence (SI) was very popular and widely used in many complex optimization problems which was collective behaviour of social systems such as honey bees (bee algorithm, BA) and birds (particle swarm optimization, PSO). This study applied a hybrid swarm algorithm namely particle bee algorithm (PBA) based on a particular intelligent behaviour of honey bee and bird swarms by integrates theirs advantages. This study compares the performance of PBA with BA and PSO for a proposed hypothetical construction engineering of TCL with materials quantity supply and demand problem. Results show that PBA performs better performance than the mentioned algorithms on a proposed hypothetical TCL problem.

Aug 1, 2013

By T. Belem

The main objective of this study is to better understand the effect of underground CPB placement conditions on its microstructural, hydro-mechanical and geochemical properties. To reach this objective, three PVC columns, each 3 m high, were built, and were filled with CPB material. The columns allow simulating undrained (UD), half-drained (HD) and fully-drained (FD) conditions along with the measurement of resulting self-weight consolidation settlement of the CPB. Bleed and drainage waters were also collected from each column for geochemical analyses (pH, Eh, EC, SO42- content). After 90-days of curing time, a total of sixty CPB samples were cored from the backfill columns and subjected to uniaxial compression tests followed by physical characterization (Gs, void ratio, water content, degree of saturation, specific surface area). The results showed that the total drainage water and the maximum observed settlement occur mainly within the first 72 hours after the columns are filled. The corresponding volumetric strains suggest that in situ backfilled stopes simulated behave in a similar way to the fully-drained (FD) or the half-drained (HD) conditions. The MIP test results show that the pore structure varies between the top and the bottom of each column.

May 1, 2007

By C. S. Maggs, L. Shwydiuk, H. W. Newcomen

The southwest wall of the Lornex Pit at the Highland Valley Copper mine has been experiencing slope displacements since 1978. Measures to manage the displacements on previous pushbacks have included stepping out, scheduling pushbacks to unload the crest of unstable areas, slope dewatering and mining with intensive slope monitoring. The final pushback on this wall is currently being mined and substantial modifications to the slope geometry could prove to be uneconomical. As a result, a campaign of surface water collection and diversion, horizontal drainhole drilling and deep well drilling is currently underway to maintain an acceptable degree of stability to the end of mine life. The various factors contributing to instability, including engineering geologic controls, rock mass strength, and surface and groundwater hydrology are presented. Movement trends, displacement magnitudes and stability analyses results are also discussed.

Jan 1, 2003

By K. Yamaguchi

"This Precious metals content copper scrap is treated with a copper smelting process, and copper and precious metals are recovered. The precious metal lost in the slag is increasing with increasing of treated scrap amount. To determine the portion of the precious metals chemically dissolved and that associated with the mechanically trapped matte in the slag will be a key factor to improve the recovery of those metals. The distribution of precious metals between copper matte and FeOx-SiO2 slag phases is determined at 1300 oC and a fixed partial pressure of SO2 of 0.1 atm for the matte grades from 40 to 70 mass% of copper. The distribution ratios show a tendency to increase when the grade of matte is increased. In addition, the solubility of platinum and rhodium in the FeOx-SiO2-CaO slag under the air and the solubility of platinum in the FeOx-SiO2 slag equilibrated with a pure platinum or the Pt-Cu alloy were determined at 1300oC in the range of oxygen partial pressure from 10-9 to 10-7. The solubility of platinum and rhodium in the slag tends to increase with increasing content of copper in the slag and oxygen partial pressure. INTRODUCTIONThe stringency of the supply and demand of platinum group metal (PGM) intensifies by localization for resources and increase of the demand for PGM, furthermore, smelting and the recycling of PGM in the nonferrous smelting process take an important role. Scrap containing a precious metal for the purpose of collection of PGM is treated by copper smelting recently. With increase of the treatment amounts, loss of PGM in slags is the serious problem.In considering of PGM loss in slags, knowledge of the distribution behavior for PGM between slag and matte phases or slag and copper phases and the PGM solubility in slags is important. The distribution ratios of silver and gold between the FeOx-SiO2 slag and the Fe-Cu matte were reported by Roghani (Roghani et al., 2000), Shinha (Shinha et al., 1985) and Takeda (Takeda et al., 1993) However, there is no data for the PGM distribution ratio. On the other hand, the PGM solubility in slags was reported by Sano (Sano et al., 1997) However, there is very few report for the solubility of PGM in the nonferrous smelting slags."

Jan 1, 2012

By Eliodor Stoian

PRESSURE information has earned a place of recognition to almost everyone interested in oil and gas production. This interest is due chiefly to the multitude of uses to which it can be put. A sub-surface pressure test, conducted for as little as $60, may yield data worth many times its cost. To the uninformed, the testing techniques may seem complex. Fundamentally, however, they consist merely of the application of common sense. At one time, sub-surface pressure data were used only in identifying individual well problems and in finding appropriate solutions. Today, it is generally acknowledged that sub-surface pres-sure .data are essential for the proper evaluation and exploitation of oil and gas pools. The following is a review of significant applications of sub-surface data with regard to: (A).-Development (B) .-Completion and work-overs (C) .-Production (D) .-Reservoir engineering (E).-Proration (F) .Other applications of pressure data of a more general nature.

Jan 1, 1958

By H. Henao, F. Canales

The present work evaluated the use of copper ore flotation tails as flux in the process of smelting copper concentrates. For this purpose, experimental and simulated liquidus lines were obtained at different ratios of tail/flux. Experimental variables included temperature, partial pressure of oxygen and sulfur dioxide and the conditions of tridymite or spinel saturation. The results indicate the feasibility of replacing the flux by increasing the temperature of the smelting process by 20°C. Results of simulation using FactSage and HSC confirm that the required increase in temperature can be achieved by increasing the flow of concentrate, enrichment of oxygen or increasing the matte grade. INTRODUCTION In the pyro-metallurgical process of copper production, the smelting stage produces two liquid phases, a sulphide phase rich in copper and another composed mainly of oxides called slag. The slag phase is the product of oxidation of the iron contained in the flotation concentrate. The smelting process which is carried out at temperatures of 1250°C, require the addition of silica as a flux and the injection of air enriched in oxygen (Schlesinger 2011). Currently, one of the most important aspects of the mining industry is the minimization of costs, the optimization of the use of available resources and the reduction of environmental impact. Under this context, the use of flotation tails from as a flux in the copper smelting process has the potential of reduced production costs and generate an environmental benefit by reduce the disposal amount of this waste material Regarding the use of copper tails as flux in the copper smelting process, a thermodynamic simulation study using the FactSage software had been reported in the literature (Parra et al., 2011). The results indicated an increase of 35°C in the temperature the liquidus for the equilibrium matte/spinel and a subsequent requirement of energy in the process of approximately 350 [MJ/ton slag]. Given that the calculated viscosity remains in the same order of magnitude, the melting temperature is a more relevant parameter for the evaluation of the use of tailings as flux.

Jan 1, 2019

By L. V. Kin, V. V. Lodeischikov, A. V. Byvaltsev, IRGIREDMET JSC, O. D. Khmelnitskaya, S. V. Petrov

"Irgiredmet Institute carried out the study on the development of gold recovery technology from refractory ore from a Russian deposit. The ore contained 1.1 g/t Au, 1.5% of pyrite and 2.4% of pyrrhotite. The recovery of gold achieved was 84.7% using gravity with generation of high grade recoverable gold concentrate and gravity middling. In the course of lime treatment, cyanidation of the middling followed by tailings detoxification the consumption of reagents was 13.0 kg/t CaO, 95 kg/t NaCN and 72 kg/t Ca(OCl)2 of the concentrate. This was capital intensive.In order to decrease the consumption of reagents, magnetic separation of the middling of the concentrates from gravity separation was added to this flow sheet. Pyrrotite recovered to magnetic fraction, which did not contain gold in accordance to this advanced flow sheet. The mass yield of pyrite goldbearing non-magnetic fraction decreased down to 1.5 times in relative to gravity middling. The slurry filtration prior to and after cyanidation allowed the organization of cyanide solution recycling.The results of pilot plant tests using the developed flow sheet showed that the content of pyrrhotite in the non-magnetic fraction decreased from 59.1% to 11.3%. The reagents consumption was 12 kg/t CaO, 3 kg/t NaCN and 16 kg/t Ca(OCl)2 of the pyrite concentrate. The total gold recovery using this flow sheet was 83.4%. Due to a decreased consumption of reagents, cost efficiency of the process was achieved. Based on the results of this research, a project design criteria was developed. A processing plant with the annual throughput of 2.6 million tonnes of ore is being designed."

Jan 1, 2016

By F. A. Auld

"Traditionally deep shaft sinking from the ground surface was carried out using drill and blast methods. Although large diameter blind shaft drilling has been carried out successfully on numerous occasions previously, including as early as 1852 in Germany (Auld, 1995), little attempt has been made until more recently to mechanise the shaft sinking excavation process using down-the-hole machines. This paper reviews the development of deep shaft sinking machines from the early 1970’s up to the present day. INTRODUCTION Mechanical excavation methods can significantly improve excavation performance, in order to access natural resources faster, and labour safety is increased compared to drill and blast operations. However, shaft sinking machines need to be capable of accommodating all the basic requirements for deep shaft sinking, namely: 1. Excavation and muck removal 2. Ground stability control 3. Groundwater ingress control 4. Temporary lining installation 5. Permanent lining installation and watertight sealing 6. Ventilation 7. Secondary means of egress for personnel In addition, how the machine operates in the shaft is an important factor, either slung on ropes from a surface headgear or on its own in the bottom of the shaft. An all-embracing machine which can deal with all of the above conditions is impossible but many of these requirements can be incorporated into a single machine. Deep shaft sinking also differs greatly from normal tunnel construction which employs the use of either a roadheader arm installed within a shield or a full-face tunnelling machine for excavation. Hydrostatic pressure and ground stability problems increase with depth, whereas in tunnelling the depth is normally relatively shallow and the aforementioned factors remain approximately constant along the tunnel length. DEEP SHAFT SINKING MACHINES All the shaft sinking carried out previously in the UK was by the traditional method of drill and blast. Elsewhere, in other parts of the world, attempts have been made to mechanise the process."

Jan 1, 2019

By A. C. T. Bigg

"The mineral potential of the Nanisivik area was discovered as far back as 1910 when a prospector of the Geological Survey of Canada, on an expedition under Captain J. G.Bernier, recorded a deposit of pyrite with minor sphalerite and galena. An outcrop of the present orebody was discovered by Texas Gulf Company in 1957. Subsequent drilling outlined the orebody consisting of 6,000,000 tonnes of ore containing 13 % Zinc and 1.5 % Lead. The orebody appears to have been formed by the filling of a limestone cavern with sulphides. The sulphide minerals consist of pyrite, sphalerite and galena. The limestone bedrock is mostly dolomite with some calcite and magnesite. A cutoff grade of about 6% Zinc has been calculated.In 1972 Mineral Resources International Limited took an option on the property. Upon completion of the feasibility study by Watts, Griffiths and McOuat, the then new independent project management group of Strathcona Mineral Services Ltd. were engaged to develop the project through the financing, construction and operating phases. Strathcona Mineral Services still manage the project from their head office in Toronto and through the Mine Manager and Department Superintendents at the Mine site. The interest for the project is presently split five ways: Mineral Resources International Ltd. 53.5 %, Government of Canada 18 %, Metallgesellschaft A.G. of Germany 11.25 %, Billiton B. V. of Holland 11.25 % and Texas Gulf Co. 6 %."

Jan 1, 1980

By R. A. Leveille, M. D. Doggett

"Economic returns for the copper industry have been assessed by building cash flow models for each of the 100 new mines brought into production over the past 20 years. Ninety-one of the 100 new mines returned the assumed 8% cost of capital and generated a positive net present value (NPV) at the development decision point. Economic criteria for individual mines are highly variable with NPV ranging from -$111 million (-$111M) to +$3850M (all money values are presented in United States dollars) and internal rate of return (IRR) ranging from -8% to +95%. The copper industry collectively would generate approximately $39 billion dollars in NPV through the development of the 100 new mines.The returns to exploration ($174M per deposit) were measured by deducting the cost of exploration per deposit ($221M) from the returns to development per deposit ($395M). Overall, the copper business has provided positive returns to exploration. However, the variability in returns across deposits means that only 41 of the 100 new mines can carry the average cost of exploration. The long lead times and high costs associated with exploration result in only those deposits with large amounts of contained copper being able to cover the average finding cost of the industry.The returns to development and exploration are shown to be highly sensitive to the cost of capital and metal price assumptions. With respect to deposit type, porphyry deposits are both larger (NPV) and exhibit lower profitability (IRR) than the nonporphyry deposits. On a geographic basis, returns to development and exploration are higher in Chile than elsewhere, reflecting the larger average deposit size and lower average exploration cost per deposit."

Jan 1, 2009

By Richard J. H. Barclay

The Lac Otelnuk iron property is located in the Province of Québec in the Northern Labrador Trough. The Labrador Trough is an extensive geosyncline of Proterozoic rocks, unconformably underlain by Archean gneisses. The Lac Otelnuk iron property hosts two significant magnetite iron-formation taconite deposits that outcrop, are sub-horizontal and are known to extend for at least 25 kilometres along strike. The host sedimentary rock package that includes a cherty iron formation is weakly metamorphosed to greenschist facies.

May 1, 2007

By J. L. Jambor, E. J. Reardon, E. C. Appleyard, J. A. Cherry

Abstract- Since the 1957 initiation of deposition of mill tailings at the Heath Steele mine, New Brunswick, periodic detailed studies of the tailings geochemistry and mineralogy have been con­ ducted, thereby providing a unique opportunity to examine the geochernical evolution of a sulfide- rich mine-tailings impoundment. When deposited, the tailings contained about 85 wt% sulfide minerals, 1 wt% to 3 wt% carbonate minerals, and 12 wt% to 14 wt% aluminosilicate minerals. Mineralogical examination of samples of the tailings from deep in the impoundment indicates that little oxidation occurred prior to tailings deposition. The mill process-water deposited with the tailings was neutral to slightly basic and contained low concentrations of Cu, Pb, Zn and As. Sampling of the tailings five years after deposition indicated that sulfide-oxidation reactions, occurring near the tailings surface, had decreased the tailings-water pH and increased the concentrations of Fe, SOi-, Cu, Zn and Pb (up to 10 000 mg/L, 25 000 mg/L, 2.5 mg/L, 200 mg/L, and 4.5 mg/L, respectively) . In 1975, ten years after decommissioning, the tailings could be divided into three zones: an upper oxidation zone in which sulfide oxidation is actively occurring, a hardpan layer, 15 cm to 20 cm thick, composed of tailings cemented by secondary Fe minerals, and an underlying reduction zone in which the tailings are relatively tinoxidized (Boorman and Watson, 1976). Chemical analyses indicated the presence of low-pH waters containing very high concentrations of SO42- , Fe, and other metals (up to 63 600 mg/L SO42-, 28 600 mg/L Fe, 10 mg/L Pb, 600 mg/L Cu and 4000 mg/L Zn), suggesting that there was a period of intense sulfide oxidation between five and ten years after tailings deposition. More recent sampling, conducted some 20 years after the end of tailings deposition, indicates that relatively little change has occurred in the last ten years, thereby suggesting that the hardpan layer has provided an impediment to the downward movement of dissolved constituents . Analyses of the tailings solids show increased accumulations of many metals at or near the depth of the hardpan layer. Evaluations of the long-term evolution of the tailings can be enhanced by comparison of descriptions of the tailings with descriptions of the gossan that overlay the massive­ sulfide orebody prior to mining. At Heath Steele, this comparison suggests that many of the mechanisms controlling the distribution of metals in the tailings will probably persist over geologic time.

Jan 1, 1992

By S. Pinard, J. MacLellan, G. Ehlert

"The Kidd Creek concentrator is looking actively for alternative technology as means of reducing operating costs in their concentrate handling section. The use of pressure and vacuum filters to eliminate the multiple hearth dryers would have the most beneficial effect. The dryers which constitute the greater portion of concentrate drying costs consume about 35,000 m3 per day of natural gas.After the evaluation of various filters by Kidd's technology group, it was clear that, because of the low final moisture requirements of 5.5 to 6% H20, few filters were able to achieve this target economically. Lately attention has focused on the Andritz hyperbaric filter. Testwork has shown that target moistures can be achieved with an appreciable reduction in operating cost.1.0 INTRODUCTIONThe Kidd Creek Concentrator is located 28 kilometers east of Timmins, Ontario. The ore is hauled by the Kidd Creek Railroad to the concentrator where it is milled to produce separate copper and zinc concentrates.Projected concentrator feed for 1991 is 3,750,000 tonnes of ore grading 3.3 % Cu and 4.7 % Zn. The concentrate production is expected to be 458,080 dry tonnes of copper concentrate at a grade of 25.7 % Cu and 265,020 dry tonnes of zinc concentrate at a grade of 54 % Zn.Currently, there is active investigation toward cost reduction through implementation of new technologies. One area which has been extensively evaluated is the concentrate dewatering and drying section. In that section, the multiple hearth concentrate dryers represent the largest source of operating costs, mainly through maintenance and energy costs (electricity and natural gas). The replacement of these dryers with filtration technology that could produce the required concentrate dryness would represent substantial savings.In September of 1990 filtration testwork carried out by Andritz Ruthner showed that Kidd Creek copper and zinc concentrates could be filtered to below 6 and 8 % moisture respectively. In late May 1991 the laboratory testwork was duplicated on siteand one month later a 5 m2 pilot filter was tested on site. Based on the pilot test results, an economic evaluation was performed. It indicated that a reasonable payback could be realized if hyperbaric filters were to replace the existing concentrate drying section."

Jan 1, 1992

By W. Dingley, G. R. Hoey

"Thirteen Canadian sulphide ore mines were surveyed for corrosion control. The mines visited were a small, yet representative, proportion of this segment of the Canadian mining and metallurgical industry. Information on metallic corrosion and methods used to control it in underground workings and mills was gathered (a) by discussing corrosion problems with company personnel, (b) by examining equipment and structures during under-ground and mill tours, (c) by photographing corrosion sites, and (d) by collecting and analyzing mine water samples.The corrosion survey, which is summarized in the re-port, leads to the following general conclusions. Mine environments are corrosive, particularly when the mine w1a-ter is acidic. Because of a lack of good corrosion control measures, excessive deterioration of equipment due to corrosion occurs in mines. Mill environments are also corrosive, but in some cases good preventive measures are being taken to effectively control corrosion losses.Ways to reduce corrosion losses in sulphide ore mines and mills are discussed. The main areas for more corrosion research and development are selection of materials of construction, protective coatings, cathodic protection, inhibitors, equipment design, and mining and milling practices."

Jan 1, 1971

By Luis A. Tondo

The Paracatu Expansion Project III in Brazil was implemented between 2006 and 2008 to increase the existing capacity from 20 to 61 Million tons per year of gold ore treatment. This expansion led the Brazilian subsidiary of the Canadian Kinross Gold Corporation to become the single largest gold producer in Brazil and the largest single gold concentrator in the world, with a nominal capacity to treat around 167,000 tons per day of ore. This paper describes the Project Development areas for the Paracatu Expansion, using the approach proposed by the Project Management Institute (PMI) and the methodology included in the PMBOK (Project Management Book of Knowledge). The author has more than 20 years of experience in the gold mining industry, having worked as a plant metallurgist, mill superintendent and development manager for RPM until he took the position as the Project Manager of the Paracatu Expansion III in 2005. Luis Tondo has a BSc in Mining Engineering from UFRGS-Brazil, an MScEng from the University of Queensland - Australia and an MBA degree from FDC-Brazil. Luis is currently based in Santiago-Chile, acting as the Project Director for developing projects in Latin America for the Kinross Gold Corporation.

May 1, 2009