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By D. A. Karvonen

In 1970, the Government of Saskatchewan instituted a program aimed at resolving the dilemma of excess productive capacity over the market demand which existed for potash produced in Saskatchewan. The two primary elements of the program were prorationing of potash production to market demand and a minimum selling price. The paper reviews the factors, both local and international in nature that led to the introduction of the program. The licensing mechanics and the operation of the program, along with the changes that have been introduced in two and one-half years of operation, are discussed. Attention is given to the results of the controls as they affect the industry and the province. Also discussed is the outlook for Saskatchewan potash in world markets.

Jan 1, 1973

By L. K. Daneshmend

"Competitive pressures dictate that the mining industry moves to larger mobile underground haulage equipment with more advanced equipment mechanization, higher levels of autonomous operation of machines, and longer periods between access by maintenance personnel. Since such new machines will of necessity be more complex than existing designs, machine reliability has become a key issue in progressing toward advanced mobile underground haulage equipment. Reliability and maintainability models of such equipment are practically non-existent in the literature. This paper addresses the issue of reliability and maintainability of mobile underground haulage equipment on two fronts: reliability modelling to enhance decision making for maintenance planning; and reliability simulation to enable evaluation of more rigorous production planning scenarios. A case study from a gold mine in the Chilean Andes is presented, for which reliability models of mobile underground haulage equipment, specifically load-hauldump vehicles and underground haul trucks, are developed, and a reliability-based simulation implemented. IntroductionMaintenance accounts for 20% to 35% of the overall operating budget for a typical underground operation, and hence represents the largest portion of the mine’s controllable operating costs (Cutifani et al., 1996; Campbell, 1997; Goddard, 1998). This, coupled with the fact that industry today is faced with an ever more complex and demanding marketplace, has led to an increasing interest in methods to reduce maintenance costs.Although significant effort has gone into developing effective maintenance strategies for industry in general, application of these to the"

Jan 1, 2003

By J. Ran

"Hangingwall sloughing is the major cause of unplanned stope dilution in open stope mining. Understanding the hangingwall sloughing mechanism can guide the stope design in minimizing the unplanned dilution and improving muck size and therefore mining efficiency. Though the modified stability graph method may be adopted to assess the hangingwall stability, for mines where the hangingwall is inaccessible, the required parameters are difficult to define and some of the simple influential factors cannot be easily identified in this method to provide guidelines for stope design.This paper identifies the main modes of hangingwall sloughing in open stope mining based on detailed data collected from mined-out stopes and site observations, applies the voussoir beam notation to analysis of the hangingwall sloughing mechanism, compares the model analysis results with the collected stope data, and provides guidelines for stope design and model application. Finally, examples are given to demonstrate the time and support effects on hangingwall sloughing and stope dilution."

Jan 1, 2002

By Omkar N. S. Beruar

Extraction of sill pillars has always been problematic because of high risk of seismicity and rockburst. Depending on a number of factors, a sill pillar can undergo different failure modes and cause problems in the mine. Of these, sudden and uncontrolled failure of part(s) of the pillar leading to highenergy seismic events and rockbursts is a main concern because it is not possible to predict the location and time of a rockburst apart from its potential to cause extensive damage to the mine structures. To date, a comprehensive technique for sill pillar design and recovery does not exist. Conventional methods of pillar design, such as the pillar stability graph method, are utilized to design stiff sill pillars. In these design approaches, the post-peak response of the pillar-rock system is not considered. This is in part due to a lack of clear and comprehensive understanding of the post-peak behaviour and failure mechanism of sill pillars. In general, the strength of a rockmass is described using a failure criterion. When stresses reach the strength of the rockmass, the rockmass does not necessarily become unstable and knowledge of the loads alone is not enough to assess the instability conditions accurately. In fact, it is not possible to predict instability of rockmass based on the strength/stress ratio alone because instability depends not only on this ratio but also on the loading condition and loading system characteristics. The loading system can dictate whether the sill pillar will go through a stable or unstable failure process. This paper examines the influence of various parameters such as the pillar width to height ratio, the extraction rate, and placement of backfill on the Loading System Stiffness (LSS) and the Pillar Stiffness (PSS). The ultimate goal is to design yielding sill pillars that would yield in a stable manner and will reduce the risk of pillar rockburst. Plane strain finite element analysis is conducted using the generalized Hoek-Brown failure criterion with elastic-brittle-plastic parameters. Examples are given to demonstrate how to calculate these stiffness values and assess the stability condition of sill pillars. It is found that increasing the rate of extraction increases the likelihood of unstable yielding of the pillar as has been found in other studies. Placement of backfill in a timely fashion in the stopes above and below the sill pillars increases the stiffness of the loading system. This indicates that backfilling of stopes while creating the sill pillar increases the likelihood of stable yielding of the sill pillar.

Aug 1, 2013

"The most widely used method for predicting the onset of continuous caving is Laubscher’s caving chart. A detailed examination of this method was undertaken which concluded that it had limitations which may impact on results, particularly when dealing with stronger rock masses that are outside current experience. These limitations relate to inadequate guidelines for adjustment factors to rock mass rating (RMR), concerns about the position on the chart of critical case history data, undocumented changes to the method and an inadequate number of data points to be confident of stability boundaries.A review was undertaken on the application and reliability of a numerical method of assessing cavability. The review highlighted a number of issues, which at this stage, make numerical continuum methods problematic for predicting cavability. This is in particular reference to sensitivity to input parameters that are difficult to determine accurately and mesh dependency.An extended version of the Mathews method for open stope design was developed as an alternative method of predicting the onset of continuous caving. A number of caving case histories were collected and analyzed and a caving boundary delineated statistically on the Mathews stability graph. The definition of the caving boundary was aided by the existence of a large and wide-ranging stability database from non-caving mines. A caving rate model was extrapolated from the extended Mathews stability graph but could only be partially validated due to a lack of reliable data."

Jan 1, 2003

By Harry Siewert, Tom Milner, Jack Stickney

"LocationThe Gibraltar mine is located in central British Columbia (NTS 93B/9W), 362 km north of Vancouver. Access is provided by 20 km of paved road from McLeese Lake, which lies 45 km north of Williams Lake on Highway 97. Latitude and longitude of the mine are 52° 30' and 122° 16', respectively.The total area of the McLeese Lake site is approximately 1700 ha (4200 acres). This includes the orebodies, dumps, surface drainage collection system, plant site, roads and tailings ponds.Brief Description of OperationOre is mined by open pit methods and is concentrated using conventional crushing, grinding and flotation processes. Copper and molybdenum flotation occurs at an average rate of 40 000 tons per day. Tailing is discharged to an impoundment area and the water recycled as process water. Additional copper is extracted from the waste dumps using sulphuric acid as a leach solution followed by solvent extraction and electrowinning.HistoryExploration by Gibraltar began in 1964, continuing until late 1970, at which time it was announced that the property would go into production. Construction started in early 1971, with initial production in March 1972. Target capacity of 30 000 tons per day was achieved by the end of March and is currently at 40 000 tons per day.Ore reserves as of January 1, 1998 were 184 384 000 tons grading 0.31% Cu and 0.010% MoS2."

Jan 1, 2000

By T. G. Joseph, G. W. Hansen

"This paper considers the duty cycle of an electric cable shovel and correlations with measured passive seismic responses of the underfoot oil sands material in-pit. It has been observed that the cyclic activity during operation of any shovel in the oil sands environment causes the material to deteriorate and provide a less than ideal footing for this large item of equipment, whose dead weight approaches 1500 tons with a footprint of merely 75 sq. yd. Consequences are manifested in the form of carbody and side frame cracks, and associated downtime, maintenance activity, cost, and subsequent loss of availability.The objective of these considerations, as part of the Oil sands-Equipment Interaction Program (OsEIP), currently a joint venture between Syncrude Canada Ltd., James Progithin International (JPi) Ltd., Caterpillar Inc., P&H MinePro and the University of Alberta, is to identify and provide feedback to the manufacturer and end user of adverse shovel-ground conditions. This would enable the manufacturer to incorporate greater knowledge into the design process and the end user to direct operations in an enhanced proactive maintenance stance."

Jan 1, 2002

By George C. Coupland

"THE PRELIMINARY EVALUATION of the orebodies was computerized for tonnages and grades on a 20- by 20-ft grid. The evaluation was based on drill-core intersections over 40-ft vertical projections coinciding with the proposed bench intervals, taking the deviations of the drill holes into account. The computer print-outs gave an ore distribution and evaluation bench by bench and indicated the areas of varying mineralization. The two orebodies have a combined length of 2,200 ft. and a maximum width of 600 ft. They strike almost due north, dip about 85 degrees east and rake to the north. The computer print-outs were used to design an open pit with an over-all waste to ore ratio of 2.58:1. At the outset, the waste to ore ratio was 4:1. The pit is tear-drop shaped, with the narrow end to the south, and at the surface is 2,600 ft long by 1,800 ft wide. Ramps are 100 ft wide, with a 10% grade. The south ramp is left-handed, starts at the southeast and will continue to the bottom of the pit. A second ramp on the west and north walls runs from surface to join the south ramp at 7 bench. This ramp will handle the total pit production when part of the south ramp, from 4 bench to surface, is lost due to the underground operation. Over-all wall slopes are 53 degrees, and are laid out to avoid prominent protuberances which could relax and become unstable and potentially dangerous (Fig. 1) ."

Jan 1, 1974

By Pierre Gauthier

Is the hydrogen economy of the future just another fish story? Absolutely not! There have been many changes affecting the energy equation over the last 10 years and conditions have changed. In this presentation, we will examine the reasons why, in the very near future, hydrogen production and distribution will be increasingly visible on many industries? radar screens.

May 1, 2007

By M. G. Lipsett

"The Mining Information Technology User Group considers information to be a strategic enabler to improved business performance in mining. This ad hoc group has been meeting since 2000 to share knowledge and leverage their efforts in effectively using IT in mining. In this paper, core issues to mining IT are reviewed, along with initiatives to engage industry professionals and decision-makers in the potential for improved use of information in their operations. Relevance of IT to Canadian MiningMining operations always drive to improve their performance, and so far they have succeeded by using different technologies. While economies of scale have come in fragmentation and material handling, and energy usage has improved through conservation and design of integrated mineral processing plants, these systems have the net result of tightening operational constraints, through reduced excess capacity. There are higher opportunity costs of lost production when integrated, large-scale systems are operating in a sub-optimal way.The problem is that it is profoundly difficult to run a mine in an optimal way. Optimization requires a defined objective function (e.g., profit), which is hard to pose, let alone solve, because so many inter-related aspects of an operation affect the bottom line. Even for a small, well-defined process state, it is difficult to measure all the pertinent information with sufficient accuracy to ensure that the system is performing well in real time.The art of mining remains making good decisions with uncertainty: in the orebody delineation; in the allocation of resources; in the efficiency of the operation; in the reliability of components and systems; and in responding to the downstream market. As well, individual decision-makers need to know how their decisions affect other processes and systems, so that improvements in one part of an operation do not result in such degraded performance in the rest of the plant that overall profitability goes down."

Jan 1, 2003

By Zhiqiang Wang

This paper mainly focuses on the method of ?three-zone? definition (a caving zone, a fissure zone and a curve subsidence zone) for full height cutting in a thick seam. Based on criteria of a caving zone and a water conducting zone, as well as existing research achievements, it is believed that there exists great error for calculation of height of the ?three-zone? area using cutting height and the coefficient of bulk increase. Dynamic factors should be considered, so, based on the theory of key rock strata and combining this with several other factors, such as mining area, this paper presents a new method to determine the overburden ?three-zone? definition. The results are as follows: (1) geometric models of a working face and key stratum are established through the rock displacement angle, with consideration of the influence of mining on the key stratum (the range of key stratum influenced by mining is enlarged as the mining area increases); (2) because there is always a cavity below the key stratum of the thick coal seam with full seam cutting, and the key stratum gradually fracture with face advancing, based on the geometric dimension of strata, the maximum fracturing interval expression has been presented, including the working face advancing range. This can theoretically lead to the location of the key stratum influenced by mining, and its stability before mining, helps to confirm which zone the key stratum and following-up strata belong to, using the criteria of sliding instability or deformation instability; (3) the new method of ?three-zone? definition which is based on the theory of key rock strata is first proposed for full height cutting in thick seam. Finally, combined with practical engineering, and comparing with field data, it is believed that the proposed method is more objective than existing methods.

Aug 1, 2013

By L. B. Hale

"Introduction This paper will discuss the current operations of Westar Mining Ltd. and give a brief history of the company and of the Crowsnest Pass area. Westar is an integrated company involved in mining, processing and marketing coal from its mines in southeastern British Columbia and shipping coal from Westshore Terminals at Roberts Bank, located just south of Vancouver, B.C. It is two-thirds owned by the B.C. Resources Investment Corporation and one-third owned by the Mitsubishi Corporation of Japan and a consortium of Japanese coal consumers. Operations in southeast British Columbia consist of four coal mines and two processing facilities (Fig. 1). The Elkview Coal Processing Plant near Sparwood, B.C. receives its raw coal from Balmer operations (Harmer open pit), Balmer South (Panel 6) an underground hydraulic mine, and Balmer North an underground room-and-pillar mine. The fourth mine, Greenhills, an open-pit operation, is located near Elkford, B.C. and has its own processing plant. The Balmer mining and processing complex is the largest metallurgical coal facility in Canada producing up to 7 .5 million tonnes of clean coal per year. The Greenhills operation, completed in mid-1983, has a capacity of 2.8 million tonnes. The high-grade metallurgical and thermal coal produced is primarily for export markets. The major customers are Japan, Korea, Brazil, Pakistan, Chile and Taiwan.Our coal and that produced by other mines in Alberta and British Columbia is shipped to the international markets through Westshore Terminals Ltd., a wholly-owned subsidiary. Westshore will have an annual capacity of 22 million tonnes of coal and accommodate ships of up to 250,000 dead weight tonnes by 1984 when its current expansion program is completed."

Jan 1, 1985

By H. L. Ramandi, P. Moore

"The corrosion of support bolts in underground mines has become an increasingly prevalent issue over the past two decades. With an increase of premature failures due to Stress Corrosion Cracking (SCC), it is evident that understanding the environmental factor influencing the cable bolt corrosion is critical for maintaining effective ground support hence the safety and productivity of underground spaces. Experimental testing of rock and cable bolts were accomplished in a variety of packings, including saturated clay, saturated coal, mine water and grout/cement to investigate their corrosion behaviours. Water samples were obtained and analysed for pH and electrical conductivity (EC) during the experiment. After a period of 76 days the bolts were retrieved and analysed visually. Primary factors were found to be the presence of geological materials and the aeration of the bolts, i.e. the vicinity to air and air flow, while secondary factors were observed to be pH levels and EC. It was also found that the coal/clay interfaces and the water/grout/cement interfaces are non-relevant to corrosion of rockbolts and cable bolts. The proposed framework can be applied to other reinforcement materials used in underground spaces for improved understanding of corrosion environment.INTRODUCTION Rock bolts and cable bolts are the major reinforcement elements used in mining industry to improve the rockmass properties for safe and productive mining. They are normally used to pin large beams or blocks of rock in underground mines to stabilise the rockmass in the roof (strata) against collapse. The rockbolts are typically low cost hot rolled rebar while cable bolts are composed of strands of wire, normally wound together around a central king wire (Hyett and Bawden, 1993; Vandermaat, 2014). The typical length of the bolts in underground mines is 2.4 m and 4–12 m for rockbolts and cable bolts, respectively (Vandermaat, 2014; Windsor and Thompson, 1997). Cable bolts are much more flexible and usually offer a much higher load capacity than rigid rockbolts. Their increased length and flexibility allows cable bolts longer than the height of a roadway to be used in underground mines."

Jan 1, 2017

By S. Klingmann, G. Blackwell

"Terrestrial photogrammetry has been used for many years to build three-dimensional pictures for analytical purposes. Photographs can include images from open pit mine walls, ground topography, mine buildings infrastructure or accident scenes. Analysis and co-ordinate estimation in early photogrammetric systems demanded the use of specialized and costly optical-mechanical stereocomparators. Few of these instruments could be interfaced with computers to provide databases, and all required trained personnel with good perspective vision to collect data.Digital photogrammetric software has replaced these older systems, and the digital video plotter (DVP) is a simple computerized system utilizing readily available PCs to quickly produce a stereoscopic model of any scene where an area of overlap exists in photographs taken from two different locations. Specialized cameras providing reference marks or grids on the photographs are still necessary for detailed analysis in three dimensions.This paper describes the application of DVP to the three-dimensional analysis of strike, dip and location of dominant fracture planes visible on an open pit wall slope, and the intersection of planar features which might reduce wall stability. Where reference survey points cannot be provided because access by personnel to the photographic scene is impossible for safety reasons, normal case terrestrial photogrammetry can provide the necessary control points for more complex aerial photogrammetry geometries.DVP recreates the geometry existing at the moment photographs are taken, providing the user with the ability to safely recreate, compare and analyze the structures determining the stability of open pit wall slopes."

Jan 1, 2003

CORPORATE SOCIAL RESPONSIBILITY CSR is about how companies manage the business processes to produce an overall positive impact on society. - Indentured work, sexual slavery, fair trade, organic growing, business ethics, environmental issues, ?Right to know -Right to decide?, AIDS activism, civil society - Balancing stakeholder issues - Government responses Who are the bad guys? Where are the threats?

May 1, 2010

By Aleksandar D. Bugarski

Despite extensive research efforts conducted in the past decades, occupational health risks associated with exposure to nano and ultrafine aerosols emitted by diesel power equipment are still not completely understood. The complexity of physical and chemical properties of diesel aerosols, constantly changing with the implementation of advanced diesel emissions control technologies, make this task rather challenging. The objective of this paper is to identify and characterize nanometer and ultrafine aerosols emitted by diesel engines in the underground environment and evaluate the effects of selected contemporary control technologies on physical properties, chemical composition, and in-vitro toxicology of diesel aerosols. The physical and chemical properties of the nanometer and ultrafine diesel aerosols are to be characterized through a test series in the NIOSH Lake Lynn Laboratory experimental mine. The crucial part of the study has been development of a facility that allows researchers to generate field results with precision and accuracy characteristics of a laboratory environment. The knowledge obtained from this study should strengthen our understanding of the health implications related to exposure to diesel particulate matter and aid in assessing the potential of various control technologies for reducing this exposure.

May 1, 2007

"Placer Dome Inc.'s Dome Mine is located in South Porcupine, Ontario and has been in production since 1910. In the past, the mill leach circuit was preceded by preaeration and primary filtration circuits. These circuits were operated to overcome the effects of pyrrhotite in the ore which reacts with cyanide to form cyanates and thiocyanates and also consumes oxygen.In early 1989, primary filters were by-passed for an extended period. Actual plant results showed no affect on gold recovery or cyanide consumption so the circuit was left off-line permanently with significant cost savings.In 1990 it was determined that considerable savings could be realized underground if backfill storage facilities were located in the mill. It was decided to convert the preaeration pachucas to backfill storage tanks on a trial basis. The trial proved successful with no effect on recovery or significant increase in reagent consumption occurring due to the elimination of the preaeration circuit. The converted pachucas worked well as backfill storage tanks.Modifications such as these have helped to keep the Dome's millings costs among the lowest of Canadian gold producers."

Jan 1, 1992

By Luc Harnois, Gabriel Voicu, Marc Bardoux, Robert Crépeau

"Abstract-Detailed petrographical and geochemical studies have been carried out on the igneous and sedimentary rocks in the Omai area, part of the Paleoproterozoic Barama-Mazaruni greenstone belt, Guyana. The stratigraphic succession at Omai begins with basalts, associated with mafic-ultramafic intrusives and poorly-sorted conglomerates; these are overlain by andesites and quartzfeldspar porphyries, with pelites and tuffaceous sediments at the top. The volcanic-sedimentary succession was intruded by a quartz-monzodioritic stock (Omai stock) and granophyric rhyolite dikes. Several generations of mafic dikes were intruded discontinuously from the Mesoproterozoic to the Permo-Triassic. Volcanic rocks range in composition from subalkaline basalt to high-silica rhyolites, with a gap between 60% to 70% SiO2, which suggests the suite is bimodal. Major, trace, and rare earth element criteria were used to subdivide the rocks into tholeiitic and calc-alkaline types. The tholeiitic rocks are characterized by flat to moderately sloping chondrite-normalized REE patterns. On the basis of immobile trace-element ratios, the tholeiitic mafic rocks from Omai have characteristics of immature intra-oceanic island arcs, or of E-MORB basalts affected by a strong arc signature, as is commonly reported for back-arc basin basalts. The andesites, quartz-feldspar porphyries, rhyolites and quartz monzodioritic stock are calc-alkaline, with strongly sloping REE patterns, enrichments in the large-ion lithophile elements, and depletions in the heavy REE and high-field-strength elements, features which are typical of more mature island-arc settings.At Omai, two types of gold-bearing veins can be distinguished: vein sets (± stockworks) and tension veins, both post-dating the regional metamorphic peak at the level they were emplaced. The relationship between the gold mineralization and its host lithologies does not seem to be genetic. Rather, the emplacement of mineralization was strongly influenced by the rheological contrasts between the intrusive/subvolcanic bodies and the volcanic/sedimentary country rocks. These contrasts helped to focus the regional stress upon the more competent intrusive subvolcanic bodies, resulting in brittle fracturing, and therefore in greatly enhanced permeability to hydrothermal fluids. The Omai gold mineralization can be classified as late tectonic, with its emplacement controlled by the last brittle to brittle-ductile stages of the Trans-Amazonian orogeny."

Jan 1, 1997

By Gabriel Voicu, Harnois Luc, Marc Bardoux, Robert Crépeau

"Detailed petrographical and geochemical studies have been carried out on the igneous and sedimentary rocks in the Omai area, part of the Paleoproterozoic Barama-Mazaruni greenstonebelt, Guyana. The stratigraphic succession at Omai begins with basalts, associated with ma?c-ultrama?c intrusives and poorly-sorted conglomerates; these are overlain by andesites and quartz-feldspar porphyries, with pelites and tuffaceous sediments at the top. The volcanic-sedimentary succession was intruded by a quartz-monzodioritic stock (Omai stock) and granophyric rhyolitedikes. Several generations of ma?c dikes were intruded discontinuously from the Mesoproterozoicto the Permo-Triassic. Volcanic rocks range in composition from subalkaline basalt to high-silicarhyolites, with a gap between 60% to 70% SiO2, which suggests the suite is bimodal. Major, trace, and rare earth element criteria were used to subdivide the rocks into tholeiitic and calc-alkaline types. The tholeiitic rocks are characterized by ?at to moderately sloping chondrite-normalized REE patterns. On the basis of immobile trace-element ratios, the tholeiitic ma?c rocks from Omai have characteristics of immature intra-oceanic island arcs, or of E-MORB basalts affected by a strong arcsignature, as is commonly reported for back-arc basin basalts. The andesites, quartz-feldspar porphyries, rhyolites and quartz monzodioritic stock are calc-alkaline, with strongly sloping REE pat-terns, enrichments in the large-ion lithophile elements, and depletions in the heavy REE and high-?eld-strength elements, features which are typical of more mature island-arc settings.At Omai, two types of gold-bearing veins can be distinguished: vein sets (± stockworks) and tension veins, both post-dating the regional metamorphic peak at the level they were emplaced. The relationship between the gold mineralization and its host lithologies does not seem to be genetic. Rather, the emplacement of mineralization was strongly in?uenced by the rheological contrasts between the intrusive/subvolcanic bodies and the volcanic/sedimentary country rocks. These con-trasts helped to focus the regional stress upon the more competent intrusive subvolcanic bodies,resulting in brittle fracturing, and therefore in greatly enhanced permeability to hydrothermal ?uids.The Omai gold mineralization can be classi?ed as late tectonic, with its emplacement controlled by the last brittle to brittle-ductile stages of the Trans-Amazonian orogeny."

Jan 1, 1997

By Xian Jian Guo, Lei Li, Hua Wang

Direct smelting of copper concentrate to blister copper has the advantages of energy saving and environmental protection. There has been a lot of effort invested in process development including test work and theoretic analysis in the past half century. However, the process has only been applied commercially to a few smelters that treat chalcocite types of copper concentrate. The main limitation to treating chalcopyrite types of copper concentrate is high copper content in smelting slag and low direct copper recovery. Thus intensive slag cleaning is needed. In this paper, the thermodynamic analysis was mainly carried out for the multiphase equilibrium of direct to blister copper smelting for both chalcopyrite and chalcocite concentrates. The practical problems in direct to blister copper smelting, which are mainly in copper content in slag, are clarified. The operational window of slag composition for different temperatures and oxygen partial pressure for the treatment of chalcocite type concentrates is identified. The process operation optimization for the blending of chalcocite, pyrite and chalcopyrite concentrates was done. INTRODUCTION The extractive metallurgy of copper can be divided into two broad groups of processes, namely, pyrometallurgy and hydrometallurgy, and the former is dominant. About 80% of copper in the world is produced by pyrometallurgy (Christopher et al., 2006). The first stage of conventional copper pyrometallurgy processing is to smelt copper concentrate to copper matte composed of FeS and Cu2S with slagging to remove a proportion of the iron (Gargul, Boryczko, & Bukowska, 2017). The smelting processes include bath and flash smelting. The second stage of the process is to convert copper matte to blister copper. In most cases, the copper matte is converted in Peirce Smith (P-S) converters, which are batch operations (Kubiszewski, 2015). The P-S converting can be divided into two periods, slagging and copper blowing. During slagging period, FeS in copper matte is oxidized to FeO, which reacts with silica flux forming slag. When the iron content in matte is less than 1%, copper begins to oxidize and reduce to form blister copper (Kucharski, Sak, Madej, Wędrychowicz, & Mróz, 2013).

Jan 1, 2019