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By David R. Mitchell, Ernest M. Spokes

There is evidence that study of minerals now considered to have susceptibilities too low for magnetic separation should be continued. Present concepts may be false. INFORMATION on magnetic properties of minerals other than magnetite is scanty and sometimes erroneous. Often there is no information at all. W. R. Crane's table of tractive forces published in 1901 is still the best guide in the field, but Crane's data can be used as a guide only, since he ignored such important factors as impurities in the minerals, size, and packing effect. During the 1930's work by the U. S. Bureau of Mines added greatly to knowledge of the behavior of certain minerals in a magnetic field. The USBM bulletin gave for the first time complete magnetic characteristics for magnetite and fragmentary data for a few other minerals. The work reported here covers a study of the magnetic-susceptibility relationships of the wolframite series of minerals and sphalerites as related to their chemical compositions. Theories are proposed to explain divergences found from commonly accepted beliefs. Gouy Method for Determining Magnetic Susceptibility: The Gouy principle has been adequately described in the literature1-3 and applied in various ways.'- It may be summarized by stating that a body of material (gas, liquid, solid, or powder mixed with gas or liquid) of cylindrical shape changes in apparent weight if suspended with one end in a zone of higher magnetic field strength than is found in that portion of the field surrounding the other end of the cylinder. This may be expressed as K2 - K1 A (H2 - H2) - mg. [1] In determining the susceptibility of mineral specimens by the Gouy method, the mineral is reduced to a fine powder and packed in a glass weighing tube. Precise measurement of the bore and depth of the tube is made unnecessary by a calibration procedure. Using the basic expression of Eq. 1, Hoare"

Jan 1, 1959

By D. H. Mullen, S. M. Runke, J. B. Cunningham

"SUMMARYThe coarse-grained granitic pegmatities of the United States have long been a source of many important industrial minerals, because of this, they have been studied intensively by geologists, mining engineers, and metallurgists. These studies have amassed a rather astounding array of published and unpublished information.Correlation of geological studies between individual pegmatites, as well as between major districts, has brought forth some semblance of order to the knowledge of the structure and Mineralogical composition of those highly complex and irregular deposits, making it possible to classify certain prgmatitites and pegmatitec zones into several groups of associated minerals; based upon the presence or absence of the minor minerals of beryllium and lithium. This classification or grouping of minerals, with the methods that have been developed for recovering certain of these minerals, makes possible a systematic approach to the winning of many of the industrially important minerals from pegmatite rock by a single integrated process.As the coarse-grained granitic pegmatites are composed primarily of,feldspar, quartz, and muscovite, the methods developed for separating and recovering them, as well as those methods developed for separating and recovering such minor minerals as spodumene and lepidolite, are important contributions to the over-all problem of pegmatite treatment. Some of these methods have been demonstrated in the laboratory only, whereas others have progressed beyond the laboratory stage arid are being employed in commercial operations.Recent laboratory investigations at Rapid City, employing both these early developments and the mineral groupings, have demonstrated that a single integrated process can recover muscovite, feldspar; and. quartz from the rejects of hand-sorting a typical muscovite-feldspar-quartz material. Employing a newly developed procedure for the flotation of beryl, the investigations have also demonstrated the possibility of recovering the beryl, in addition to muscovite, feldspar, and quartz, by an integrated process from a typical berylliferous pegmatitic material in which all of these minerals are present."

Dec 1, 1952

By T. Holmes

Investigations reveal that the throughput difference between a perfect production day and the average is often on the order of 40%, and that bulk solids handling system deficiencies are the largest (>50%) class of lost opportunities. Root causes for underperformance of elements in material handling systems can often be traced back to the detailed design and even study-stage oversights. To reduce or eliminate bulk solids handling problems in an existing system requires an analysis of troublesome handling equipment – for example, a concentrate transfer chute that experiences plugging, poor flow to the furnace, unacceptable dust generation, and spillage. These problems often lead to significant maintenance costs related to downtime and personnel time. Understanding the cause and cost of flow issues is the first step in troubleshooting the problem. Next, measuring the material’s flow properties is critical to providing design data required to develop a comprehensive scientific solution. In the case of a new installation, the design data is required to properly design the system right first time and prevent flow problems from occurring so that the new equipment works correctly from start-up. This paper highlights the cost of material handling problems and discusses the best approach to evaluate material handling system deficiencies. A case study is given of how Anglo American's operation in Barro Alto Brazil enjoyed a 5-month payback from a bulk solid transfer chute debottleneck project. INTRODUCTION When operating a plant or designing a greenfield or brownfield operation, the intent is to have perfect production days (PPD) each and every day. Money is made by moving material through the plant in a controlled manner to produce the desired product. Mining operations and copper concentration processes are at the most basic level a practice in bulk solids handling (BSH). Unfortunately, many of the project planning steps are overlooked, particularly related to BSH. Investigations reveal that the throughput difference between a PPD and the actual average is often on the order of 40%, and that BSH system deficiencies are the largest (>50%) class of lost opportunities (Wellwood, 2017). Root causes for underperformance of elements in material handling systems can often be traced back to the detailed design and even study-stage oversights. BSH issues are essentially baked into the fabricated plant and left to operations to accept the deficiencies and deal with them for the life of the mine. These deficiencies then become part of the normal operation of the mine and plant.

Jan 1, 2019

By William M. Knuth

This study involved a review of bonding and bond release practices in the states of the Eastern and Midwestern Coal Provinces. Data was collected from regulatory authorities and site visits to document the bonding and release procedures in the states related to achieving successful reclamation. The regulations related to PL95-87 were reviewed. The analysis included suggested modified or new bond release criteria, a review of applicable bond release inspection techniques was prepared, a conceptual graduated bonding system and incentive procedures for reclamation of surface mine sites to alternative or higher land uses. A number of conclusions and recommendations for further research were included. A detailed bibliography and a glossary are included in the report.

Jan 1, 1978

By Kazutoshi Haga, Batnasan Altansukh, Atsushi Shibayama

In this study, the dissolution behavior of copper and arsenic from five different copper concentrates containing 21-34.7 %wt. of Cu, 1.3-7.3 %wt. of As and 15.6-27.5 %wt. of Fe were investigated under high pressure oxidative leaching conditions. Leaching experiments were conducted using different leaching reagents such as ferric sulfate, sodium chloride, deionized water, sulfuric acid and ferric sulfate-sodium chloride mixture in order to compare the effectiveness of each on the dissolution of copper and arsenic from copper concentrates. The effects of ferric sulfate (0.1-1 mol/L), sodium chloride (0.1-0.5 mol/L) and leaching temperature (100-180C) were investigated under 1 MPa autoclave pressure , particle size >45 m, pulp density of 100 g/L, stirring speed of 750 rpm and leaching time of 60 min. Results showed that the most effective leaching reagents for the dissolution of copper are 0.34 M NaCl and a mixture of 0.1 M Fe2(SO4)3 - 0.34 M NaCl, in which As extraction did not exceed 7%. The copper concentration in the pregnant leach solution obtained after leaching using 0.1 M Fe2(SO4)3 - 0.34 M NaCl was 13-30 g/L, while the concentration of As and Fe were in a range between 0.3-1 g/L and 0.2-10 g/L at 160C. INTRODUCTION Copper sulfide minerals, such as chalcopyrite (CuFeS2), chalcocite (Cu2S) and covellite (CuS) are used for the production of copper in mining and metallurgical industries worldwide. The conventional pyrometallurgical processing followed by electrorefining is a common process for the production of high purity copper (Moskalyk, & Alfantazi, 2003; Padilla, Aracena, & Ruiz, 2012). However, the degradation of copper content and an increase in the amount of metal impurities, especially arsenic (As), antimony (Sb) and bismuth (Bi) in copper concentrates cause serious problems to the production of copper and environmental pollution (Basha, Selvi, Ramasamy, & Chellammal, 2008; Long, Peng, & Bradshaw, 2012). Hydrometallurgical processes are used to recover metals from refractory ores, concentrates and tailings by leaching (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018). During the last decades, hydrometallurgical processing for the recovery of copper from enargite-bearing copper ores and concentrates has been extensively investigated due to its relatively more environmentally benign and less capital costs than the conventional pyrometallurgical processes (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018; Choubey, Lee, Kim, & Kim 2018). However, several issues facing the hydrometallurgical processing are incomplete extraction of copper associated with the coprecipitation of iron compounds, difficult stabilization of toxic impurities and production of a contaminated copper product (Dreisinger, 2003; Safarzadeh, Moats, & Miller, 2014). Hence, a number of studies have been conducted on the selective dissolution of copper from copper ores and concentrates or selective removal of arsenic leaving a clean copper residue using various leaching reagents or lixiviants (Safarzadeh, Moats, & Miller, 2014; Mihajlovic, Strbac, Zivkovic, Kovacevic, & Stehernic, 2007; Tongamp, Takasaki, & Shibayama, 2009). More extensive studies on the selective extraction of copper from enargite- bearing copper ores and concentrates are much needed to improve copper recovery.

Jan 1, 2019

By J. N. Van der Merwe

The current state of the South African mining industry is briefly reviewed from the point of view of an engineer not schooled in the finer details of predictive economics. Based on the assumption that current world exploration targets are indicative of which commodities the world will want in the future, it is concluded that the South African mining industry will continue to be viable and growing well into the future. South Africa holds the minerals the world wants and needs, and as long as the world has people, the demand for those, including gold, will continue to grow. There still is a growing demand for vanity commodities' and even the demand for platinum as a jewellery material displays significant growth. Mine safety still needs to improve, but significant improvement has already been made although the industry is not always acknowledged for the effort in this field. The safety target set for 2013 is within reach but will require sustained effort. The physical mining conditions for the main revenue earners, coal, platinum, and gold, will change substantially in the near future. Research in mining has declined alarmingly and it is urgently necessary for that to be revitalized to generate the knowledge that will be required. In order to streamline the transfer of knowledge to the mining engineers of the future, research needs to be conducted with the full involvement of the universities. It will also be necessary for the universities to continually review curricula to incorporate the latest knowledge and to embark on a campaign of knowledge transfer to the older generation of mining engineers. The mining industry is well supported by the universities, who have done their duty with regard to societal transformation, and the universities are in turn supported by the industry through vehicles like the METE Industry is also strongly supported by a number of professional and vocational societies. The SAIMM will continue to play its vital role as the meeting point of the mining technical sciences, and will continue to disseminate knowledge via the journal, schools, and conferences. In the future, it will review its modus operandi to support the needs of growing membership in the remote branches. It will enter the research arena in an appropriate manner and give effect to the developing integration of mining and society. The most important threat to the future of the mining industry is the recurrent mention of nationalization. While interaction between the mining industry and political and economic development of South Africa has always been characteristic, the current debate is emotional more than rational, and engineers are not well equipped to participate. This matter needs to be finalized before it becomes a popular force that politicians cannot control.

Jan 1, 2011

These may be of interest to members examining the collection of ores, etc., from Mount Lyell, which is a feature of the Victorian Gold Jubilee Exhibition, and may have some value as geological records. No. 7 and No 8 are analyses of samples of footwall rock from the Lyell Tharsis mine. No. 7 is hard, crystalline, fine-grained, and of homogenous structure. It has a dark-red colour, due to ferric oxide (Fig 1); and was taken from a place eight feet east of the ore body and about half way between the 50 feet and 100 feet levels. No. 8 is from the same excavation and from four feet east of the ore body it is softer than No. 7, being easily scratched with a knife-blade, of a deeper dark-red colour, and has a slaty cleavage.

Jan 1, 1902

By Sara J. Davenport

The human race has been exposed to copper for thousands of years, as copper and gold were the metals first used by man (1). if According to Ramazzini (2), "of all the metals that man's ingenuity has contrived to dig out of the bowels of the earth, iron and copper are the most useful, far more so than silver and gold." Workings of copper in the oldest known copper mine associated with the smelting of this metal have been traced back to 4300 B. C. Facts revealed by archeological research in India, Austria, Hungary, France, Italy, and Spain indicate that the use of copper probably began about 5000 B.C. (3), and that about 4000 B.C. a copper age came intermediate between the stone and bronze ages (4). Copper ores are rather widely distributed and are found in nearly every country. Many of the copper minas of Europe have been worked since before the Roman Empire. The must important mines in Europe occur in the district around Huelva, extending into Spain and Portugal. The German deposits are located mainly about Mansfeld (Rhenish Prussia); they have been worked continuously for 1,000 years. Some of the mines in Africa have been worked by the natives for centuries, and China has produced copper from shallow pits for thousands of years. India also has many ancient mines worked in the early centuries (2).

Jan 1, 1953

By X Zhang, B McFadzean, C T. O’Connor, S Pikinini

The Platreef in South Africa contains one of the world’s largest reserves of platinum group elements. One of the major platinum group minerals (PGM) components in this ore is sperrylite (PtAs₂), which has been found to be slow floating compared to the other PGMs. In a study aimed at determining the reasons for the poor floatability of sperrylite and to improve its recovery, tests were carried out to determine the interactions of the mineral with a range of standard and novel collectors. Microflotation tests were carried out to determine the hydrophobicity of the minerals before and after adsorption of collectors. In addition, experiments were carried out to determine the enthalpy and extent of collector adsorption on the minerals and to investigate electrochemical interactions between sperrylite and selected standard collectors. The results indicated that poor collector adsorption was the major cause for the poor floatability of sperrylite. Sperrylite had both poor natural (14.5 per cent) floatability and collector-induced flotation recoveries, the highest recovery of 26 per cent being observed using PNBX collector at pH 9. Moreover, the rest potential tests showed that the interaction of sperrylite with standard collectors such as xanthates was poor. The decrease in potential after the addition of the standard collectors ranged from 7.9 mV to 24.9 mV under alkaline conditions. Furthermore, sperrylite rest potentials remained unchanged with a change in pH from basic to acidic, indicating its resistance to surface alteration. Surface alteration of the minerals is key for the chemisorption of collectors to take place on the mineral surface. This was further emphasised by flotation recoveries at low pH, where only certain collectors responded positively to a decrease in pH and these were shown to have undergone a structural change. This work has highlighted the difficulties associated with sperrylite flotation and provides some direction for future research.

Aug 24, 2022

By M. Dobbins

The advent of the more efficient and effective rare-earth roll (RER) separator in the early 1980?s began a new era of dry magnetic separation. The RER units quickly became the separator of choice within the materials processing industry. Though RER separators are now commonplace throughout the world, only modest advancements have been made to the original design. This paper details recent, more significant, rare-earth magnetic separation developments. These developments maintain the best features of existing technology while minimizing shortcomings. The discussion includes discerning features that aid in better metallurgical performance. Also discussed are improved mechanical aspects that allow for both easier operation and maintenance. Performance details from various test programs are included.

Jan 1, 2009

horizontal-type classifier prior to charging into a blastfurnace.

Apr 1, 1995

By J B. Magee

Sheep Creek Mines Limited has developed several operating techniques that have greatly assisted in keeping unit costs at reasonable levels. Some of the techniques, as described in this paper, include organization, accounting, warehousing, bonus and contracts. INTRODUCTION THIS paper is a non-technical report on operating techniques developed by Sheep Creek Mines Limited at their Mineral King mine, located near lnvermere, B.C. Depressed metal prices and continually rising labour and materials costs have necessitated a continuous search for method to keep operating costs down. The newly developed techniques and practices have greatly assisted in maintaining unit costs. Sheep Creek Mines Limited is a mining company incorporated under the laws of the Province of British Columbia. The Company's share capital is 3,500,000 common shares, witl1 a par value of 50 cents each, of which there are 1,875 000 shares is-sued. The Company has eight di-rectors. The administration, exploration and accounting offices are located in Nelson, B.C. The Company's working capital is approximately $1,500,000.

Jan 1, 1963

By H. B. Charmbury, D. R. Mitchell

Everyone in the coal industry from top management to the preparation engineer is vitally interterested in recovering as much salable coal as possible from the run-of-mine product. Coal losses from a preparation plant occur in the solid refuse material being discharged from the plant and in the bleed-off slurry from the water circuit. The refuse is generally hauled from the plant in solid form by trucks, conveyors, or aerial trams to the refuse or culm banks or it may be crushed and transported by hydraulic methods to refuse ponds or dams. The loss of coal in refuse material and the extent of this problem has been covered in another paver.' The loss of salable coal in the plant bleed-off slurry is one of the water circuit problems in coal preparation plants. Slurries originate at various places in the processing circuits and are usually disposed of by transportation to waste ponds known as slurry ponds. Even though the combustible material discharged as slurry and refuse from the plant is labeled a loss, the loss must only be considered temporary. There is a tendency at the present time to examine refuse and slurry losses and to reprocess these materials to recover valuable coal. Slurries from eight bituminous coal preparation plants in central and western Pennsylvania have been sampled and examined to determine the amount of recoverable coal in each. Examinations of the samples were made to determine the tons per hour of solids and the gallons per minute of water in the slurry; the size, ash, and sulfur analyses of the solids; and the sink-float characteristics of the larger solids. The extreme fines (—325 mesh material) from one of the samples has been examined further to determine the micron size fractions as well as the ash and sulfur content of the micron sizes. Collection of the Samples: Field trips were made to each of the plants where samples were obtained and discussions held with the preparation engineer prior to sampling. The flowsheet of the plant operation was studied and a time-period selected to obtain a representative sample. At least 20 1-gal samples of total plant bleed-off suspensions were collected at each plant in increments over designated periods belived to be typical of the plant flow Three of the samples were obtained from plants in the Pittsburgh area. Each plant supplied information on the capacity, operating time, type of cleaning equipment used, water and material balances, and auxiliary equipment. From the time required to collect given amounts of slurry and from the amount of solids in the slurry, total solids discharges from the plant were calculated and compared with the plant flow. Ash and sulfur analyses of the total solids were also checked, at some plants, with company figures. In all cases where checks could be made there was reasonable agreement between company figures and those reported here. Analysis of the Slurries: Representative fractions of properly agitated samples were examined in the laboratory to determine the specific gravities of the slurries. The solids and water were separated by filtration to check the weight of each so that the percent of solids in the pulp could be determined. From the resulting data, presented in Table I, it may be seen that the percent solids in the slurries ranges from 2.8 to 39.0 and the amount of solids from 2.61 to 45.61 tph. Size Analysis of Solids: Representative fractions of the slurries were permitted to settle and the clear water was removed by siphoning. The solids were then wet-screened at 28, 35, 48, 65, 100, 150, 200 and 325 mesh. Results of the size analyses are presented in Table 11. These results indicate the extreme variation of sizes and in most cases the degree of fineness of the solids. The amount of —325 mesh ranges from 22.5 to 84.5 pct and the amount of+200 mesh from 8.3 to 69.1 pct.

Jan 1, 1961

1. Scope of project In South Australia, the Western Mining Corporation operates the underground copper and uranium mine Olympic Dam. The deposits concerned are enormous (approx. 570 million tons) and have been mined since 1988. At the end of 1996, Western Mining decided to increase the capacity of Olympic Dam from 85,000 t.p.a. to 200,000 t.p.a. Planning At this point, SIEMAG-GHH Mining Division was asked to carry out a study including the planning of the entire shaft hoisting operations and calculation of the investment and operating costs involved. With the assistance of Australian partners, the study was completed within a single month. The principal points of the study were as follows: ? Planning of the hoisting machinery for 1,375 tons per hour from the surge bin below ground to the discharge bunker above ground. ? Optimisation of the hoisting operation with regard to maximum availability and safety. ? Optimisation of the time schedule to minimise the financing costs of the investment. ? Statement of the investment and operating costs. ? Comparison of the technical data, design and availability of the planned machinery with reference plants. Construction of the plant Following acceptance of the study, the decision was taken within a few weeks to implement it. SIEMAG Transplan GmbH received the order to execute the entire shaft hoisting machinery on the basis of the study submitted. The decisive factors in placing the order were: ? Convincing study, adherence to schedule ? Reference to comparable plants elsewhere ? Reference with regard to availability.

May 1, 2003

By B. Matsetela, S. Amos, V. Nkuna

"The Kamoa-Kakula copper project—a joint venture between Ivanhoe Mines, Zijin Mining Group Co. Ltd, and the government of the Democratic Republic of Congo—has been independently ranked as the world's largest undeveloped high-grade copper discovery by international mining consultant Wood Mackenzie. It is a very large, near-surface stratiform copper deposit with adjacent prospective exploration areas within the Central African Copperbelt, approximately 25 km west of the town of Kolwezi and about 270 km west of the provincial capital of Lubumbashi. A phased development approach is currently envisaged. The first phase of the project involves the development of a 6 Mt/a underground mine and surface processing complex (dual stream concentrator) at the Kakula deposit, a discovery announced in early 2016. The second phase involves an integrated 12 Mt/a two-stage development, beginning with initial production from the Kakula mine, to be followed by a subsequent, separate underground mining operation at the nearby Kansoko mine, along with the construction of a smelter. Project description and locationThe Kamoa–Kakula project comprises a newly discovered, very large stratiform copper deposit with adjacent prospective exploration areas, located within the Central African Copperbelt in Lualaba Province, Democratic Republic of Congo (DRC). The Kamoa–Kakula project lies approximately 25 km west of the town of Kolwezi and about 270 km west of Lubumbashi. Ivanhoe owns a 49.5% share interest in Kamoa Holding, an Ivanhoe subsidiary that currently owns 80% of the project. Zijin Mining owns a 49.5% share interest in Kamoa Holding, which it acquired from Ivanhoe in December 2015 for an aggregate cash consideration of US$412 million. The remaining 1% interest in Kamoa Holding is held by privately-owned Crystal River.The Kamoa–Kakula project consists of the Kamoa exploitation licences (exploitation permits 12873, 13025, and 13026, which cover an area of 397.4 km2) and one exploration licence (exploration permit 703 covering an area of 12.74 km2). The Kamoa exploitation licences, approved 20 August 2012, grant Kamoa Copper the right to explore for, develop, and exploit copper and other minerals, for an initial 30-year term, expiring on 19 August 2042. The permits can then be extended for 15-year periods until the end of the mine’s life."

Nov 1, 2018

By Lawrence Devon Smith

"The concept of risk is well established in the mining industry. It is acknowledged when estimates of reserves are expressed as proven or probable, when mining and metallurgical recoveries are applied to the ore, and when contingencies are added to costs. However, it is difficult to provide a quantitative assessment of risk. The significant sources of risk in a mineral project and how these are addressed in project evaluations are discussed. The discount rate is examined as a fundamental means of reflecting risk in discounted cash flow evaluations. Current industry practice is discussed, and a methodology for the analysis of risk levels is proposed that assesses the constituent components of the discount rate: real interest, mineral project risks, and country risk.IntroductionThis paper examines the components of the discounted cash flow discount rate and proposes a method of estimating project-specific discount rates. The motivation for this work comes from experience of project evaluations in which owners and purchasers had agreed on virtually every aspect of the evaluation except how much the project was worth. Agreement was reached on reserves, grad~, recovery, capital costs, operating costs, taxes, and, by combining these components, even on the fmal cash flow values. The only difference in opinion concerned the discount rate to be used in the calculation of the net present value. Depending on the life of the project, such differences of opinion can cause a variation of more than 50% in the value placed on a project! This is illustrated in Figure 1."

Jan 1, 2003

By O. C. Blade

Studies on more efficient utilization of petroleum have been made for 50 years by the Bureau of Mines. As part of this work, the Bureau has analyzed thousands of samples of crude oil by its standardized routine method, and many of these analyses have been published, sometimes in special compilations, other times in connection with reports of research on petroleum. This report is a finding index for all analyses published from 1921 to 1959, and brings up to date the information in Information Circular 7470 3 and in the appendix of Bulletin 490.4 The routine method of analysis and the equipment originally were described in a report by Dean and associates. 5 This report has long been out of print, but Bulletin 490, bringing the procedure and its modifications to date, was published in 1951. 6 Briefly, the procedure consists of testing, by standard methods, small samples of a crude oil for gravity, sulfur, color, viscosity, pour point, and (more recently) nitrogen; and, by a standardized procedure, distilling a portion of the crude oil and determining certain properties on the separated fractions and residuum. Careful records are made during the procedure, and the distillation data obtained are used to compute the approximate amounts of products present in the sample--such as percentages of gasoline, kerosine, gas oil, and lubricants. The principal data and results are recorded on a standard analytical report form. An analysis of a sample of crude oil from a recently developed field is shown in table 1. Compilations of these analyses, sometimes in tabular form and usually for specific geographic areas, are published from time to time. Analyses often are included in connection with other studies, such as engineering field reports.

Jan 1, 1959

By Andreas Jungbauer

The Chinese steel producer Rizhao Steel is now focusing on the highly attractive local and export markets for high-quality, thin-gauge strip products. The first line of five Arvedi ESP plants began operations in the spring of 2015. The new casting-rolling complex is designed for a total annual production capacity of 11 million tons of highquality, ultra-thin, hot-rolled strip products with widths of up to 1,600 mm and thicknesses down to 0.8 mm. The energy consumption and related costs are reduced by up to 45% compared with conventional casting and rolling processes. This also means a major reduction in CO2 emissions. During the project phase, the plant personnel from China received extensive training at the licensor’s plant – Acciaieria Arvedi S.p.A. in Italy – to prepare them to deliver the best operational performance with the new state-of-the-art endless casting-rolling facility. The plant setup, the training and support program, as well as the latest results from the start-up are presented in this paper.

Aug 16, 2017

By B. R. Broekman, K. J. R. Ford

Over the last decade significant development and use has been made of a simulation model that describes the HCFeMn smelting operations in Assmang Ltd. This spreadsheet based, semi-empirical, predictive model is derived from data from the submerged arc furnaces at Assmang?s Cato Ridge, KwaZulu-Natal smelter site. Departures from the empirical base case mass balance due to changes in the ores, reductant feeds and operating parameters are accounted for by use of fundamental relationships and heuristic type ratios which are applied in an iterative manner in the model. The model performs a sequential set of calculations to iteratively converge on a fully consistent mass and energy balance solution, which predicts Mn alloy and slag output masses and compositions for a given set of feeds. 440 MWh per operating day is used as the basis of energy input. The paper describes the key model components, assumptions made and calculation approach. Added to the mass and energy balance is a cost and revenue balance, which allows prediction of the economic performance for each simulation. Substantial use of the model has been made in simulating the performance of the current Assmang HCFeMn furnaces for both smelter investigations and future production planning. The model has also been a valuable tool in the strategic analysis of the use of future Mn ores and has formed the vital link between mine ore grades (metal and slag elements) and the expected alloy output, grade and profitability. Additionally, the model has been used in benchmarking exercises to predict other Mn smelting operations with remarkable accuracy. The paper highlights some aspects of the approaches taken in using the model to benchmark Assmang?s own furnace performance and cost competitiveness with other producers. For Assmang the development and application of this modelling technology has led to transformation in understanding of the Assmang HCFeMn furnace operations and increased awareness of the potential of its Mn ore bodies.

Jan 1, 2004

By D. L. Paulson

The Bureau of Mines is investigating pyrometallurgical and chemical methods for extracting copper from sulfide ores under conditions that provide for control of gaseous and solid emissions. Under this program, a new electrothermic reduction process was developed to produce blister copper after collecting essentially all of the sulfur by a single-unit operation. The process feasibility was indicated by a laboratory-scale investigation that demonstrated chalcopyrite concentrate can be roasted to less than 1 percent sulfur and that the resulting calcine can be reduced with carbon to blister copper by a cyclic process in a single arc-smelting furnace. Pilot-scale tests in an 800-kV-A electric-arc furnace demonstrated that 98.4 pct of the contained copper can be continuously recovered as metal by operating at bath temperatures between 1,350° and 1,375° C. Smelting rates of approximately 225 lb of blend ft2 of bath surface/hr resulted in power consumptions of approximately 520 kW-hr/ton of blend. The blister copper that was produced contained less than 0.1 pct iron while the precious metals reporting to the copper product and gangue material elimination were similar to those in a conventional smelting operation.

Jan 1, 1976