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By V. K. Srivastava1

Coal mine fires burn about 20?21 million tons of coal per year throughout the world, which not only results in loss of valuable coal reserves, but also cause serious environmental hazards over large areas. The famous Jharia Coalfield of Dhanbad, India is no exception; here a coal mine fire originally reported nearly a century ago is still continuing unabated in vast areas of the coalfield. In this paper, an attempt has been made to analyse high resolution multispectral thermal images of ASTER for mapping and delineating surface and subsurface coal mine fire areas of this coalfield. From the generated temperature gradient map, as well as from the surface temperature anomaly map, it has been possible to differentiate shallow and deep coal mine fire regions and other erroneous cultural features. It is observed that the coal fires are distributed mostly in the eastern part of the area, which has branched into two major coal fire zones: one along Lodna- Tisra-Bhulanbarari-Kujama-Jealgora Jharia and the other along the Kusunda-Kendudih area. The western fire zone is restricted in Dumra-Nadkhurkee-Jayramdih area. From ASTER thermal infrared data from 2006, the total affected area is estimated to be about 7 km2, of which the deep subsurface coal mine fire area is around 3 km2 and the shallow surface coal mine fire area is around 4 km2, as compared to 18 km2 coal mine fire area estimated in 1987 using Landsat TM thermal band data. In other words, the fireaffected region has decreased by about 50 % during the last 18 years. However, such decrease in coal fire area is seen in the central portion only, and has in fact increased in the south eastern parts of the Jharia coalfield. Study of Differential Interferometery images, generated from Earth Resources Satellite (ERS) synthetic aperture radar (SAR) tandem data has shown that the active coal mine fires are in general associated with low topography/subsidence areas.

Aug 1, 2013

By L. Beiza, L. Velásquez-Yévenes

Chalcopyrite is by far the most abundant copper sulphide on Earth, but it has proved to be problematic in its processing by leaching, mainly related to its refractory nature, slow kinetic rates of dissolution and the formation of an inhibiting product layer on its surface during leaching. This study is focussed on the electrochemistry of chalcopyrite dissolution in acidic chloride system, which has recently gained attention as a feasible alternative to conventional sulphate media for chalcopyrite leaching. It is well understood that chalcopyrite dissolution from mineral grains depends strongly on the solution redox potential at the mineral surface as well as the composition of the product layer that may form on its surface. The chloride ions in solution support the Cu(II)/Cu(I) couple as the primary oxidant, as they can form complexes with the cuprous ions formed, increasing the dissolution rate. Additionally, these ions increase the potential range where chalcopyrite can be oxidized. The chemical reaction that occurs on the surface of chalcopyrite in chloride solutions in the presence of cupric ions will be characterized and the coulometry of the reaction will be reported to understand what is happening on the surface of chalcopyrite. Electrochemical experiments have been carried out to investigate the behaviour of mixed potential by open circuit potential (OCP) measurement for 30 minutes using sulphuric acid at pH 1 under nitrogen at 25⁰ C. The effect of chloride was analysed at 0, 20, 50 and 100 g/L and cupric ions from 0 to 50 g/L. The potentials obtained were then used to carry out chronoamperometry tests for a period of 30 minutes in absence of cupric ions. The mixed potential did not show a significant difference with an increase in cupric concentrations in the absence of chloride. Increased mixed potentials and current densities were observed at higher chloride and cupric concentrations. Chronoamperometry tests at those potentials were used to construct a voltammogram trace which has the same trend as a normal cyclic voltammogram but with much lower current densities. INTRODUCTION As a result of the depletion of copper oxides, the leaching of copper sulphide minerals has become the focus of many studies. Chalcopyrite, (CuFeS2) which is the most abundant of copper minerals, has proved to be challenging to process via hydrometallurgy, mainly related to its refractory nature, slow kinetic rates of dissolution and the formation of an inhibiting product layer on its surface, during leaching.

Jan 1, 2019

By A. Sargeant

This paper evaluates the application of post consumer glass as a cementing agent in underground mine backfills. The underlying theory indicates that glass is pozzolanic and, when used in a finely divided form, behaves as an activator to generate binding products, thus contributing to the development of backfill strength. The objective of this study was to evaluate the strength performance of backfills when Normal Portland Cement is replaced with various percentages of post consumer glass. The study evaluated the performance of backfills prepared using tailings from three mines. Each backfill was prepared using unique recipes, and were evaluated at cure times of 7, 14, 28, 56, 112 and 224 days. Laboratory work involved visual, physical and chemical material characterization followed by strength evaluations. Results of strength assessment revealed that backfill prepared with glass is reactive and can contribute to the strength of backfill. The results also indicated that the reactivity of glass may be affected by the incorporation of slag within the recipe, the availability of lime and the level of hydration of the binders. The study concluded that, at 15 % glass replacement of NPC, backfill prepared with mine tailings, NPC and glass can achieve improved or equivalent strength performance compared to backfill prepared tailings and NPC alone.

May 1, 2007

By I. Silin, M. Stapelmann, H. Wotruba, K. M. Hahn, C. Gerold, J. Katzmarzyk

New deposits are often characterized as low-grade and complex ore bodies. To date, processing of fine grained ore deposits implies complex processing routes with high energy consumption. For that reason, the investigation of dry grinding, e.g. with vertical roller mills (VRM), is an approach to save energy reserves. It is already proven that dry grinding can influence subsequent processing steps. In terms of flotation, the surface characteristics of the minerals are crucial for a successful flotation process, e.g. the sufficient adhesion of mineral particles and air bubbles. Wet grinding results in different surface properties of the product entering the conditioning stage than dry grinding, which has direct influences on the subsequent flotation process. To investigate the consequences for flotation, processing of a copper sulfide ore containing chalcopyrite and bornite was investigated, including grinding and rougher flotation. The effects of wet tumbling mill and dry VRM grinding on the degree of liberation of valuable minerals are characterized using mineral liberation analysis by scanning electron microscopy. The obtained flotation results, as well as the recorded flotation parameters such as dissolved oxygen, oxidation reduction potential and conductivity are presented and discussed. INTRODUCTION Copper production is mainly connected to porphyry copper deposits as main source for copper sulfides such as chalcopyrite and bornite. State-of-the-art processing of these deposits is already established worldwide for decades, combining crushing, fine grinding and subsequent flotation with rougher, cleaner and scavenger stages (del Villar, Desbiens, & Maldonado, Bouchard, 2010). While the demand for copper is increasing, copper grades of the deposits are decreasing and the remaining ore bodies have more complex mineralogy and finer particle size distribution. This implies processing routes with high water and – due to the liberation of fine grained valuable minerals – high energy consumption by fine grinding. One of the main energy consumers in processing plants is the grinding section, which accounts for more than 50% of the energy consumption in processing plants and even 4% of the worldwide energy consumption (Jeswiet, & Szekeres, 2016). Therefore, dry fine grinding alternatives are applied to enhance the overall energy consumption of mineral processing plants: high pressure grinding rolls (HPGR) and vertical roller mills (VRM).

Jan 1, 2019

By P. Latostenmaa, T. Veneranta, J. Järvi, E. J. Peuraniemi

Boliden Harjavalta has operated a copper flash smelter since 1949 thus the smelter celebrates 70th anniversary of flash smelting in 2019. This paper shortly describes the historical perspective of the smelter and gives an update of the current operational issues and process modifications made during the recent years. The plant has carried out several major investments lately providing a firm view on the future. A new landfill area for the final slag was commissioned in 2018. A modernization of the concentrate storage area, a new acid plant and a debottlenecking project of the smelting line have been recently commissioned. At the same time, the tankhouse capacity will be expanded. This paper focuses on smelter modernizations. INTRODUCTION Boliden is a metal company operating mainly in northern Europe with core competencies in the fields of exploration, mining, smelting and metal recycling. Boliden Harjavalta, located in Finland, is part of the Smelters Business area of the Boliden Group and operates a copper and nickel smelter with related sulphuric acid plant in Harjavalta, and a copper refinery in Pori. As a short historical review, Harjavalta’s history may be dated back to the year 1910 when the Outokumpu ore body in eastern Finland was discovered. By the end of 1913, a copper mine was in operation and a copper plant based on Hybinette process was ready for commissioning. Due to difficulties in metal recoveries and throughput the Hybinette process was permanently shut down in 1929, and 100% of the mining production was sold as concentrate abroad. As the mine production increased steadily, a decision of a new smelter was made in early 30’s. In 1936 Outokumpu’s copper smelter based on electric smelting was commissioned in Imatra. The smelter was dismantled and transferred to Harjavalta during the final stages of the Second World War. The copper refinery started in Pori already in 1941 and the first copper anodes were cast in Harjavalta in 1945. First sulphuric acid plant started operation in 1947. Due to the post-war shortage of electrical energy, the plant metallurgists initiated a major development work to start exploiting the latent heat of concentrates. Thus, the flash smelting method was developed in Harjavalta and commissioned in 1949. Ten years later, flash smelting was also applied for nickel concentrates. Table 1 summarizes major historical and technical milestones of Harjavalta. The history of flash smelting is described in details in another paper of this conference (Jåfs et al., 2019).

Jan 1, 2019

By Adrian White

Agenda Overview of pastefill at Kidd Mine Pipeline wear and the need to monitor Steel Loop monitoring Steel Borehole monitoring Monitoring pipe linings

May 1, 2013

By M. Thérien, M. Rémillard

"Data from two industrial plants are treated by mass balance package BILMAT. The first application concerned a base metals/precious metals (PM) producer and the second, implementation of a hydrocyclone for the recovery of iron chips in the granite industry. Also LOTUS 123 software can easily be adapted for laboratory test mass balances using an IBM-PC computer. INTRODUCTIONThe present paper deals with two different BILMAT program applications. The first application concerns a copper-zinc-gol d-sil Yer producer based in North-western Quebec and the second, concerns the iriplementation of a hydrocyclone for the recovery of iron chips in Quebec granite company.Afterwards, we deal with some LOTUS 123 software applications for laboratory test mass balances using an IBM-PC.1. Base metal and P.M. producerNoranda Inc, Mattagami division, which operates mines in North Western Quebec, produces copper and zinc concentrates from a sulphide ore. Small amounts of gold (0,34 g/t) and silver (20 g/tl are present. At the time of the study, only 30% of the silver content was recovered in the copper concentrate. The main objective of this study is, to evaluate the metallurgical behaviour of silver in the flotation circuits, to increase its recovery in the copper concentrate and therefore to improve its final economic value. It is then worthwhile to identify the silver bearing minerals and to determine the critical flotation stages where they are lost. The methodology proposed is based on the coupling of a mineralogical study and a detailed mass balance ca1cu1ation in the flotation circuit performed using the BILMAT mass balance package (1, 2, 3).The results are discussed with respect to the distribution of silver in the circuit. The recoveries obtained at the various separation stages for the minerals present in the ore are also discussed."

Jan 1, 1987

By Stavros A. Argyropoulos, Radhakrishna Valiveti, Bernard Closset

"The use of microprocessors in metallurgy is currently increasing at a very fast rate. A large number of companies are offering systems ""off the shelf"" which are aimed at the data acquisition and process control market. There are two main drawbacks to these systems: first, they are usually quite expensive, and second, they are not completely compatible with the target metallurgical application. An alternative is to develop a microcomputer-based system "" in lab "".A dedicated microcomputer system has proved to be very useful in many metallurgical research applications(1,2)). However, in this system the analog input signals had to be transferred over long distances , and significant noise was introduced despite the use of shielded cables. In order to overcome this problem a network of low pass-filter s was designed and built. The usage of the abovementioned system brought the need to digitize the analog signals in-situ and transfer the digital signals, which are much less susceptible to noise over very long distances. This distinctly different design philosophy has become a reality with the evolution of microprocessor technology over the last few years. This design philosophy is utilized in this development."

Jan 1, 1984

By W. Almdal

"To compete in the ever expanding global market, companies must continuously improve. Benchmarking, a tool that can help determine where the best areas for improvement are, is more than comparing measures, it is comparing how business is carried out and how decisions are made. The key to success is in integrating benchmarking results into the organization, so that those most affected by the results own them and are empowered to use them as a positive change agent. This is a necessary step to achieving substantive change to practices and achieving optimum performance improvement. IntroductionBenchmarking is a powerful tool that can assist in the continuous goal of improvement. Benchmarking is more than comparing productivity rates or unit costs, it is seeking out best practices and assimilating them into a business to provide a positive change agent.Benchmarking is a continuous process - is time consuming, labour intensive, and requires a lot of discipline. There are no easy answers and no easy solutions, just a lot of work. The results, however, are nothing short of spectacular. More and more companies are adopting benchmarking practices because it is a key tool of the learning organization and will become part of the way business is done, part of the way it is managed. Business must change to stay ahead or to get ahead and benchmarking can accelerate the process of business change. Improvements in an organization can happen faster and more effectively with a benchmarking approach."

Jan 1, 1994

By R. F. Dawson, A. S. Bagnall, K. Barron

"At Smoky River Coal Limited (SRCL), economic occurrences of structurally thickened coal deposits form attractive surface mining targets. These steeply dipping, structurally complex deposits require that special attention be given to pit wall design and operational practices.Where anchor support is feasible, there is potential for reducing footwall waste requirements and/or increasing coal recovery. SRCL has used rock anchors to support footwall slopes since 1985. This paper outlines design considerations and performance of rock anchor support systems at SRCL.Three case histories are included that illustrate the design process for supporting steeply dipping footwall slopes using rock anchors.IntroductionWestern Canada contains vast reserves of high grade coking coal largely located in two belts within the Rocky Mountain Foothills straddling the Alberta/British Columbia border. There are about one billion tonnes of ""clean"" coking coal reserves (Romaniuk and Naidu, 1986) situated near existing mining infrastructure. Within the reserve areas shown in Figure I, a total of six producers operate 10 large mines linked with ports situated on Canada's west coast. Al! of the producers operate conventional shovel/truck surface pits.Surface mineable reserves within some of the mountain coal mines are located on dip slopes or where the geologic structure has thickened the coal near the surface, creating low strip ratio targets. Footwall slopes in these mines are developed parallel to bedding, thus minimizing waste extraction and preventing undercutting of potentially unstable strata. The base of the mineable coal measures in both north and south regions is marked by massive quartz sandstone units (Torrens Member and Moose Mountain Member) which often form the footwall. Footwalls developed within this basal sandstone unit can attain very high unbenched heights often with little risk of instability. In cases where the footwall is developed on weaker strata above this basal unit, stability considerations control slope height when bedding dips exceed about 30°. In these situations it is desirable to assess the benefits of artificial support to increase the unbenched footwall slope height."

Jan 1, 1995

By M. J. Brodie

"The decision to proceed with mine decommissioning, generally involves a number of complex and independent factors. These factors, including potential for re-opening, legal requirements, contingent liability, asset depreciation, due diligence, and public relations are discussed in the context of the financial implications on a net-present- value basis. It is concluded that the decision to delay mine decommissioning is commonly not in the best financial interests of the mine operator. Some measures to reduce decommissioning costs and expedite the process are presented.IntroductionThe primary audience for this paper is mining executives, although it should be of interest to all professionals involved in the planning and carrying out of mine decommissioning. Historically, mines were abandoned, not decommissioned. Then, as decommissioning became required, it was assumed that salvage revenue would pay for decommissioning. During the I980s, it became apparent that mine decommissioning costs would exceed salvage revenue for nearly all sites, and that a substantial net cost for decommissioning would be realized for some sites. Throughout this period, the approach to decommissioning has evolved through the sequential addition of measures to meet the increasing regulatory burden. This has resulted in a process in which the elements of decommissioning may not be optimized with respect to each other for the most favourable net present value to the corporation. Additional considerations which are not specific to mine decommissioning, such as environmental policy, due diligence and public relations, may now come to bear on the decision-making process.This paper addresses the issue of mine decommissioning as a three-step process. The first step involves identification and evaluation of all of the issues that relate to mine closure, the most importantof which is the potential for re-opening of a mine. Assuming that it is decided that a mine is unlikely to re-open, the second step is to determine when to carry out the decommissioning work. This aspect is examined through a net present value (NPV) approach. Finally, details regarding who does the work and how the various tasks are divided up must be resolved.Every mine is a unique concentration of minerals at some specific location, hence the development, operation and closure of a given mine must be regarded as a special case. Superimposed on the site-specific considerations are the size, history and financial status of the associated corporation."

Jan 1, 1995

By R. F. Dawson, A. S. Bagnall, K. Barron

"At Smoky River Coal Limited (SRCL), economic occurrences of structurally thickened coal deposits form attractive surface mining targets. These steeply dipping, structurally complex deposits require that special attention be given to pit wall design and operational practices.Where anchor support is feasible, there is potential for reducing footwall waste requirements and/or increasing coal recovery. SRCL has used rock anchors to support footwall slopes since 1985. This paper outlines design considerations and performance of rock anchor support systems at SRCL.Three case histories are included that illustrate the design process for supporting steeply dipping footwall slopes using rock anchors.IntroductionWestern Canada contains vast reserves of high grade coking coal largely located in two belts within the Rocky Mountain Foothills straddling the Alberta/British Columbia border. There are about one billion tonnes of ""clean"" coking coal reserves (Romaniuk and Naidu, 1986) situated near existing mining infrastructure. Within the reserve areas shown in Figure I, a total of six producers operate 10 large mines linked with ports situated on Canada's west coast. Al! of the producers operate conventional shovel truck surface pits.Surface mineable reserves within some of the mountain coal mines are located on dip slopes or where the geologic structure has thickened the coal near the surface, creating low strip ratio targets. Footwall slopes in these mines are developed parallel to bedding, thus minimizing waste extraction and preventing undercutting of potentially unstable strata. The base of the mineable coal measures in both north and south regions is marked by massive quartz sandstone units (Torrens Member and Moose Mountain Member) which often form the footwall. Footwalls developed within this basal sandstone unit can attain very high unbenched heights often with little risk of instability. In cases where the footwall is developed on weaker strata above this basal unit, stability considerations control slope height when bedding dips exceed about 30°. In these situations it is desirable to assess the benefits of artificial support to increase the unbenched footwall slope height."

Jan 1, 1995

By A. de Ville d’Avray, G. Guezennec, P. Wavrer, E. Braak

The aim of the European CEReS project was to develop an innovative copper recycling route based on the coprocessing of coal production waste by bioleaching and printed circuit boards (PCBs) from Wastes of Electrical and Electronic Equipment (WEEE). Coal waste is processed by bioleaching in stirred tank reactors, and the resulting bio-lixiviant solution is used to extract copper from PCBs pre-treated by catalytic cracking. Copper is then extracted using solvent extraction, stripping and electrowinning. The whole process flowsheet combines thermal, biotechnological, physical and hydrometallurgical operations. Once the process concept has been validated at laboratory scale, a simulator of the whole process has been built in the USIM PAC software platform. The tool enables to improve the analysis of the results of the lab-scale experiments and to design industrial-scale scenarios by simulation. The results of such simulations allow to estimate the economical interest and to perform environmental impact assessment. After presentation of the CEReS project objectives and main results, the paper describes the methodology used to build the simulator and define the scenarios corresponding to feasible process routes. The main results of the simulations are then presented with technical, economic and environmental point of views. INTRODUCTION Because of rapid technological innovations in the electronics industry, electronic waste represents the fastest growing waste stream in the world, especially in developing countries (Işildar et al., 2018). Meanwhile, the global copper demand is expected to go on rising in the future. As significant amounts of copper are present in end-of-life printed circuit boards (PCBs), it becomes a necessity to develop efficient copper recycling routes. Metal refining from e-waste is traditionally performed by pyrometallurgical processes. However, the hydrometallurgical processing of PCBs, which has been shown to be technically feasible at laboratory scale with a wide range of leaching reagents, could be a less energy intensive and more flexible alternative (Tuncuk et al., 2012). However, a large amount of leaching reagent is required to leach the metals from PCBs.

Jan 1, 2019

By Sean Dessureault

Technological Revolution ?Information age ? brought about by: ?Low-cost processing, storage, information sharing ?Shared & integrated data ? Increasingly more sensors & data sources From data collection to data use ??Data mining??analyzing the huge volumes of data ?Key example used in MBAs: ?Wal-mart, Google Explosion of data, trickle of information ?Paralysis of action Outline Trying to provoke- What drives innovation now? ?Challenges in mining Data Action Approach ?Examples from research & consulting upon request Conclusions ?Unaddressed challenges (Research Opportunities)

May 1, 2006

By D. G. Macintyre

Thick westward prograding clastic wedges were deposited along the margin of ancestral North America from Mid-Proterozoic to Mid-Paleozoic time. Periodic rifting and tectonic subsidence resulted in the formation of starved sub-basins along this otherwise passive continental margin. These sub-basins host important sedimentary exhalative barite-zinc-lead-silver deposits. In northeastern British Columbia, a thick succession of Paleozoic basinal facies clastic rocks is preserved within the northwest-trending Kechika Trough. Within the trough, sedimentary exhalative barite and barite-sulfide deposits occur in starved basin sediments of Middle Ordovician, Early Silurian, and Late Devonian age. The last are the most economically sigrrificant and are hosted by carbonaceous cherty argillites and siliceous shales of the Middle to Late Devonian Gunsteel Formation of the Lower Earn Group. The largest deposit, Cirque, has reserves in excess of 35 million tonnes averaging 100J0 (Pb +Zn) and 47 g/t Ag. Other potentially significant deposits include Driftpile Creek, Bear, Mt. Alcock, Fluke, Pie and Elf. Stratiform barite deposits are also common, particularly near the basin-to-shelf transition zone. The Late Devonian barite-sulfide deposits typically are zoned: interlarninated barite-sphalerite-galena-pyrite occurring near suspected vent areas grades outward into bedded barite. Laminated beds of pyrite typically occur in hangingwall siliceous shales, particularly above the inferred vent zone. The apparent absence of epigenetic stringer sulfide zones and footwall alteration implies that the deposits were formed from ponded brines at relatively low temperature. Away from the main deposits, the favourable stratigraphic interval is marked by the presence of thin beds of nodular and laminated barite and thin laminae of pyrite. Thin tuff beds also occur locally in this stratigraphic interval and are evidence for a Late Devonian volcanic event. A paleogeographic reconstruction of the Kechika Trough in Late Devonian to Mississippian time suggests that the barite-sulfide deposits formed in third-order starved basins between uplifted, northwest-trending horst blocks that were capped by shallow-water carbonate banks. The presence of intraformational breccias suggests that fault movement accompanied barite-sulfide precipitation. This tectonic activity may have allowed the escape of heated, overpressured metalliferous brines from permeable reservoirs within the sedimentary pile. Sedimentary exhalative deposits formed where these brines were exhaled as buoyant plumes onto the seafloor, cooled, and ponded in anoxic seafloor depressions.

Jan 1, 1992

By J. Robitaille, P. Cousin

"Agnico-Eagle Mines Ltd, Laronde Division, is a gold operation processing ore containing copper and zinc. Copper is recovered as a concentrate prior to gold cyanidation.Copper concentrate grade was improved with the implementation of an on-stream analyzer. Further improvements to the copper concentrate grade could be achieved throughefficient zinc depression.High cyanide cost linked to the copper content in the flotation circuit tails, prompted some modifications to the flotation circuit. Economics and environmental issues were considered. A Maxwell cell was retrofitted with an Outokumpu HG cell mechanism stabilizing it's operation. Secondly a flotation column was added to increase circuit capacity.This paper will address flotation circuit operation difficulties, successes and results achieved before and after modifications.RESUMELes Mines Agnico-Eagle division Laronde sont considéré comme étant un producteur d'or. Le minerai révèle toutefois des teneurs appréciables en cuivre et zinc. Un circuit de traitement combinant des étapes de flottation du cuivre et de cyanuration de l'or a été adopte. Le rejet de la flottation est lixivie pour son contenu en or.La teneur moyenne en cuivre du concentre final fut augmentée suite a l'implantation d'un système d'analyse en continu. Une séparation plus efficace de la sphalérite se devra d'être obtenue afin d'améliorer davantage la qualité du concentre produit.Le contenu en cuivre au rejet du circuit de flottation influence directement la consommation en cyanure. Pour cette raison, certaines modifications à la flottation sont devenues nécessaires. Des facteurs d'ordre économique et environnemental ont été considérées. Une cellule existante fut transformée en cellule HG d'Outokumpu, stabilisant son opération, et une colonne de flottation fut ajoutée au circuit afin d'accroitre sa capacité.Le présent document se veut une rétrospective des problèmes d'opération vécus, des solutions apportées ainsi que des résultats obtenus et de leurs implications suite aux améliorationseffectuees."

Jan 1, 1995

By K. M. Tait, M. M. McDonald

"An overview of the status of the development and application of Integrated Gasification Combined Cycle (IOCC) power generation is presented. Current projects in North America, Europe and Japan are described, together with the status of plans to apply the technology in Canada. A brief review of the technology is presented, including promising coproduction cycles and hybrid gasification fluid bed processes. The ability of the process to be modified for CO2 to be extracted for enhanced oil recovery or disposal is addressed. The environmental impact of IGCC technology is compared to those of other power generation technologies IntroductionDefinition and HistoryIntegrated Gasification Combined Cycle (IGCC) technology is a concept which has been around for about 50 years but has been generating increasing interest in the last 10 to 15 years. This paper focusses on coal as the feedstock for gasification. Other feedstocks used in gasification systems have included oil, biomass, tires and petroleum coke. IGCC technology integrates coal gasification with generation of electricity by means of a combined cycle system which fires the produced gas in the gas turbine. Figure 1 shows the basic components of an IGCC system which are the gasifier, the gas cleanup system and sulphur recovery plant, the gas turbine/generator set, the heat recovery steam generator (HRSG) and the steam turbine/generator set. The benefits of an IGCC system over a conventional Rankine Cycle pulverized coal (RCPC) fired system for the generation of electricity from coal are that the IGCC system has potential to have higher efficiency and lower emissions of NO., SO and CO2 per unit of power produced than the corresponding RCPC system."

Jan 1, 1994

By Robert B. Parsons

"Two recent decisions by the Canadian courts could reduce the taxes paid by many profitable mines in Canada. At a time when the mining sector is hurting from an ever increasing tax burden and soft prices, the results of two cases are welcome news. Mining companies are well advised to check prior years' tax returns in the context of these developments. IntroductionSince 1976, the income tax legislation has provided a reduced rate of tax for taxpayers engaged in the mining of mineral resources, through the mechanism of the ""resource allowance"". The deduction given for the resource allowance effectively reduces the federal rate of tax by one quarter on resource profits. In general terms, ""resource profits"" are defined to include a taxpayer's income from: the production and processing in Canada of ore to any stage that is not beyond the prime metal stage (or the pellet stage or its equivalent in the case of iron ore); and, royalties computed with reference to the amount of value of production from mining operations in Canada.In computing income from these sources, the taxpayer must deduct expenses that may reasonably be regarded as applicable to those sources, with the exception of expenditures for Canadian exploration and development or financing costs.As a result of the mechanics of the calculation of the resource allowance, one quickly realizes that the lower the expenses deducted from the resource profits, the higher the resource profits and consequently the higher the 25% resource allowance. The impact of the recent court cases is to narrow the range of expenses which are required to be deducted in the determination of the resource profits.The two cases are outlined below."

Jan 1, 1991

By R. D. Lindsay

Originally the focus of Deak Resources Corporation Business Plan was to establish a method for treating small gold and possibly base metal mineral deposits located in the Kirkland Lake - Noranda area of Northeastern Ontario, Northwestern Quebec at a central treatment facility. Since the establishment of this original business plan in 1988 other processing options have availed themselves for handling the base metal ores off site at a central facility which may preclude their processing at Deak.The site selected for the Deak processing plant was Kerr Addison's Virginiatown operation. At the time this property was acquired from the Receiver following Golden Shield Resources' Bankruptcy in 1989 the mill had been processing in the order of 1, 100 SDT per calendar day through one processing module; a module originally designed to process 1,500 SDT/day of a free milling gold ore. This tonnage level however had not been achieved since 1966 when an average throughput of 1481 SDT per calendar day was recorded.In order to realize Deak's overall Business Plan it would be necessary to expand the mills processing capabilities to provide:(1) A 2200 SOT per day circuit capable of processing a free milling, medium work index gold ore;(2) a 1 ,000 SDT per day flotation and/or cyanidation circuit capable of processing refractory gold bearing materials containing refractory minerals (i.e. copper) that should be removed prior to cyanidation; and possibly(3) a 1200 to 1500 SDT/day base metal circuit capable of processing the broad range of Deaks' Base Metal Property ores. If this circuit were introduced it would be completed in two tonnage phases.Further, it was recognized when operations resumed under Deak in 1990 that the existing tailings dam was suitable for, and had sufficient capacity for additional gold tailings with necessary physical modifications and changes in disposal techniques (including the construction of containment storage cells) that would allow storage for an additional 3.4 million tons of gold ore tailings approximately 1.0 million of which has been deposited since 1990. The focus of Deak's long term tailings disposal strategy is to acquire the Barber Lake Basin which is currently adjoined to our main tailings system which will provide an additional 10 million tons of storage capacity.

Jan 1, 1993

By William Petruk, Donald Leroux, Michael Cooper

"Brunswick's primary flotation circuit (CuPb) produces a low grade concentrate containing considerable amounts of undesirable sphalerite and pyrite. To determine the cause of the poor selectivity, a series of Quantitative Mineralogical Analyses (QMA) were performed on several CuPb circuit streams. It was observed that significant fractions of sphalerite and pyrite reporting to the concentrate was liberated, meaning that their depression was not complete. The study highlighted the need for some R&D efforts to be spent in sulphide flotation fundamentals.IntroductionThe viability of many mining projects strongly depends on the performance of their concentrators. More than ever, competitive grade concentrate and high metal recoveries are criteria of paramount economic importance. The selectivity of solid/solid separation processes used in concentrators must be maintained at high levels or continuously improved.Flotation is the solid/solid separation process used by most base metal producers. In spite of its low operating cost, flotation remains a very complex process. To make it selective, precise reagent dosages and electrochemical conditions must be maintained. The difficulty in controlling flotation lies in the numerous heterogeneous reactions that can take place between the different solid, liquid and gaseous phases involved. This complex chemistry makes flotation selectivity difficult to understand and, therefore, to improve from a theoretical point of view.Brunswick Mining and Smelting, near Bathurst N.B., operates a 10500 tpd concentrator which produces four concentrates through five stages of sequential flotation (Figure I). Various reagents are added at key process stages while the ore undergoes different electrochemical conditions. Although the selectivity is important at all stages, the performance of the very first separation, the CuPb circuit, is the most critical. It is here where mineral misplacement to the wrong stream has the greatest economic consequences."

Jan 1, 1996