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By G. Voiloshnikov, P. Koblova, Ye. Musin

Preg-robbing activity of the carbonaceous matter negatively impacts the process of precious metals recovery from gold-bearing ores and concentrates. In this paper, ore heap leaching (HL) using a blinding agent which inhibits ‘preg-robbing’ is a promising method. A carbonaceous ‘preg-robbing’ gold ore with head gold grade of 1.92 g/t from a Russian deposit was studied. Corg was 1.12%. In accordance with the Irgiredmet technique, preg-robbing activity was 65.3%. Surfactant reagent DN was used as the blinding agent. Column tests on the ore HL were conducted using two options. They were: a conventional flow sheet (Column 1) and a flow sheet with the reagent addition (Column 2). Gold recovery in Column 1 was 55.39% while gold recovery in Column 2 was 68.48%. The test results confirmed the efficiency of reagent DN. Keywords: Carbonaceous ore, organic blanking, preg-robbing, recovery

Jan 1, 2020

By Y. Wang

Arsenic-bearing acidic effluent from hydrometallurgical processes contains many harmful metal ions and must be appropriately treated before discharge. In the present study, arsenic, copper, zinc, aluminum, and magnesium were co-precipitated by means of the Fenton reaction. The precipitates obtained under different conditions were investigated to determine their stability. The results indicate that pH value and hydrogen peroxide (H2O2) dosage have significant effects on the removal of various elements. Arsenic, copper, zinc, and aluminum (but not magnesium) can be removed at pH 5–6 and anH2O2/As mole ratio of 2 at ambient temperature. The precipitates were mainly amorphous and granular with particle size in the micrometre range. The arsenic concentration in leachate from the toxicity characteristic leaching test was 3.6 mg/L, which proves that the precipitates are effective in fixing arsenic.

Aug 3, 2023

By D. Orynbassar, N. Madani

This work addresses the problem of geostatistical simulation of cross-correlated variables by factorization approaches in the case when the sampling pattern is unequal. A solution is presented, based on a Co-Gibbs sampler algorithm, by which the missing values can be imputed. In this algorithm, a heterotopic simple cokriging approach is introduced to take into account the cross-dependency of the undersampled variable with the secondary variable that is more available over the entire region. A real gold deposit is employed to test the algorithm. The imputation results are compared with other Gibbs sampler techniques for which simple cokriging and simple kriging are used. The results show that heterotopic simple cokriging outperforms the other two techniques. The imputed values are then employed for the purpose of resource estimation by using principal component analysis (PCA) as a factorization technique, and the output compared with traditional factorization approaches where the heterotopic part of the data is removed. Comparison of the results of these two techniques shows that the latter leads to substantial losses of important information in the case of an unequal sampling pattern, while the former is capable of reproducing better recovery functions.

Aug 1, 2021

By L. Zhou, J. Qiu, M. Tang, J. Wang, G. Dai

Coal spontaneous combustion in longwall gobs is a major concern in coal mines. In comparison to other hazards, this phenomenon is difficult to detect. Fifteen factors influencing coal spontaneous combustion in longwall gobs were identified from the literature. The logical relationships between these factors were evaluated and quantified by decision-making trial and evaluation laboratory (DEMATEL) and interpretive structural modelling (ISM) methods. A four-level hierarchical structure model was established. The research shows that combining DEMATEL and ISM can reveal the key factors that influence coal spontaneous combustion in longwall gobs and establishthe interactions and relationships among these factors. The results can provide the basis for preventing and controlling coal spontaneous combustion in longwall gobs.

Jun 10, 2024

"Chairman: Professor C. O. Frush Rapporteur: Mr A. C. SchmidtPapers:Planning of underground copper mining by J. K. Williams, L. Smith and M. P. WellsLinear programming models of copper leach and solvent extraction plants by W. P. C. Duyvesteyn, J. R. Cutland and J. R. L. RusselPlanning and forecasting stope production by D. W. Kappes and P. F. X. Moussett-JonesA proposed method of dynamic ore reserve assessment for a caving system of mining by A. F. Spinks and S. P. NicholsThe initial point made by Mr L. Smith in presenting his paper was that the planning aid outlined in the preprint was neither sophisticated nor complicated. However, it was a system which worked and was a very real way in which the computer was being used to assist miners, in particular, who at the present moment have very little experience in the use of computers.Mr Smith apologised for the fact that the editors of the paper had eliminated all reference to 'time'. This was very unfortunate because the essence of the system was that it saved time, which was one of the main points made in the paper.The operations being planned by this means cover the Rokana orebody which is a stratiform body mined by open stoping.The planner has to decide how much ore to mine from each base in each time period of his plan, that is, from three to four months hence to a year ahead. Normal procedure is for the planner to have discussions with the underground manager on some of the information he requires. The sort of information needed is (i) tramming limits for each face, (ii) extraction and grade factors, (iii) the present status of mining operations so that the planner knows where to start the new plan, and, (iv) details of any other mining consider¬ations that need to be incorporated in the plan.The objective at this point is to produce an initial plan for discussion purposes. All the information required comes from one sheet per stops and shows, inter alia, the stope identifier, the length along strike, the ore tonnage, the copper and cobalt grades and, finally, the extraction and grade factors.The program itself, among many other things, checks the data for validity, compatibility, completeness and credibility and prints error messages accordingly. Output from this system is in the form of a summary of production from each face, as well as giving the figures of ore reserve depletion. This summary indicates which slopes are to be mined in this time period, the original survey grades (copper and cobalt), the estimated production grades, how much is to be mined from each stope and the copper and cobalt production accordingly."

Jan 1, 2014

In many practical situations, the samples available to estimate a block of ore are not located on a regular grid. The sample sizes and orientations are not constant, and the number and relative position of the samples used for valuation of each block change from block to block. Estimation of a block W might be very complex and time consuming if each sample in the neighbourhood of the block were considered individually. A practical solution consists of grouping the samples in blocks Wi (i = 1, 2, . . . n) in the neighbourhood of W, and estimating µw as a weighted average of the mean xi of the samples in Wi. Possible ways of grouping the samples are illustrated in Fig. 1 1.1. More complex patterns can be found in the literature (Krige, 1966) To calculate the kriging estimator of pw, we must be able to calculate the variance of xi, the covariance of xi and xj, and the covariance of xi and pµw. The purpose of this chapter is to show how these quantities can be calculated when we ignore the position of the samples in the block Wi. We consider the samples to be randomly distributed in Wi; kriging with this assumption is known as random kriging.

Jan 1, 1978

By S. R. Grobler, S. Bilessuris, M. E. Greenway

Ensuring that all the head ropes on a multi-rope friction winder share the load equally has been a design and maintenance issue since they were first developed. Accurate machining of the rope grooves is necessary to meet this objective. Early analysis of the change of rope loads for ropes in grooves that are mismatched in depth treated the grooves as rigid. Groove depth tolerances deduced from this analysis are stringent and difficult to achieve in practice. Test work indicated that the inherent flexibility of groove lining materials alleviates the load-sharing problem. However, no analysis of rope load changes due to mismatched flexible grooves has been published. In this paper we develop new equations for the rope load variation for flexible grooves. By allowing flexibility to reduce to zero, the corresponding rigid rope groove equations are obtained. New criteria for tolerable groove depth variations are developed from this analysis, which depend on the flexibility of the groove lining material.

Sep 1, 2021

By G. P. O’Malley, A. N. Nikoloski

One of the most promising options for extraction of copper from low-grade chalcopyrite ores by heap leaching, however, chalcopyrite is refractory and suffers slow leaching kinetics and poor overall extraction due to formation of a passivation layer on the particles. The solution composition is believed to strongly affect the formation of this layer. This paper examines the effects of the major cationic species of Fe3+/2+, Mg2+, Cu2+, and Al3+ and the anionic SO42− species in acidic ferric sulfate media with a composition simulating recycle heap leach conditions. It was found that altering the sulfate, iron or aluminium concentration had significant effect on the copper extraction, the concentration of copper slightly influenced the extraction, and the amount of magnesium had no appreciable effect. Also, iron and aluminium caused a change in the sulfate concentration, and the sulfate ions had the greatest influence on the extraction of copper under recycle leach conditions

Jan 1, 2023

By A. Drumond, J. F. C. L. Costa, M. G. Lemos, A. L. Rodrigues, F. G. Niquini

The past decade has seen a renewed possibility of using machine learning algorithms to solve a large collection of problems in several fields. Data acquisition for mining operations has increased with the growth in sensor-based technologies, and therefore the amount of information available for mining applications has dramatically increased. Ore sorting equipment is available for separating ore from waste based on differences in physical properties detected by a real-time analyser. The separation efficiency depends on the contrast in these properties. In this study we investigate the application of machine learning models trained using data from the output of a dual-energy X-ray ore sorting apparatus at a gold mine. The particles were first hand-sorted into ore and gangue classes based on their mineralogical composition. Classification models were then used to help decide the balance between the number of true and false positives for ore in the concentrate, with a view to economic parameters, using their receiver operator characteristic (ROC) curves. The results showed AUC (area under the ROC curve) scores of up to 0.85 for the classification models and a maximum reward condition Fpr/Tpr around 0.5/0.9 for a simplified economic model.

Aug 8, 2024

By E. Wroth, D. G. Krige

Written contribution E. Wroth* (Visitor): The technical aspects of the assistance scheme which IS provided for in the Gold Mines Assistance Act, 1968, and certain consequential amendments to the Income Tax Act, 1962, are explained by Dr Krige. The object of this contribution is to explain the working of the Gold Mines Assistance Act (hereafter referred to as the 'Assistance Act') for the benefit of those concerned with the tax and assistance assessments of gold mining companies. The views expressed do not necessarily reflect the Revenue Department's interpretation of the legal provisions involved. 1. MINES WHICH QUALIFY FOR ASSISTANCE 1.1 The assistance scheme is applied on a selective basis. Mines which qualify for assistance are defined in the Assistance Act as 'assisted gold mines'. The requirements for classification as an assisted gold mine are that: (i) the mine be a producing gold mine; (ii) it would be likely to close down within eight years if not assisted; and (iii) with State assistance its life would be appreciably prolonged and there would be a significant increase in production of gold or uranium, or of both gold and uranium. 1.2 A mine is only considered for classification as an assisted gold mine on application. Such applications must be made to the Mining Leases Board through the Government Mining Engineer. Classification of a mine as an assisted gold mine is at the discretion of the Minister of Mines who acts in consultation with the Minister of Finance and has regard to the recommendations of the Mining Leases Board. 1.3 The Minister of Mines will impose conditions subject to which a mine is to be classified as an assisted gold mine. The most important will be that the mine must immediately lower its operating pay limit, as normally calculated, by a percentage fixed by the Mining Leases Board and conduct its future mining operations in accordance with the average grade of its ore reserves determined on the basis of such lower pay limit. l.4 The Minister will fix the date as from which classification as an assisted gold mine is to be effective. So long as the mine has operated on the basis of the lowered pay limit for at least three months during the particular year of assessment, he would normally fix the effective date as the first day of the year of assessment.

Jan 5, 1968

By M. Royle, A. Bloem, R. Uken

Rock mass damage induced by mining increases the fracture frequency and joint aperture width in the pit walls and rock mass behind open pit mine slopes. The depth and intensity of this mechanical damage relative to the pre-mining state is not well understood and yet is a major factor in the development of slope stability models and operational guidelines. The induced damage can be partitioned into domains or zones. The blast damage zone is essentially a ‘free flow’ rock mass volume, which transitions quite rapidly into the excavation damage zone, which extends beyond that. In both zones, transient high in situ water pressures are often linked to unstable slope conditions. Whereas rock damage can enhance passive depressurisation in the rock mass, high infiltration rates on the other hand, as one could expect during rainstorm events, will decrease slope stability within the dilated pit rock mass, especially in the blast damage zone-excavation damage zone, by increasing pore pressures for short periods of time. Ice jacking and freshet are additional factors to consider in (sub) arctic environments and can be exacerbated by unfavorable rock fabric orientation. On a local scale, brittle deformation damage zones, in the form of faults, can contribute to slope instability, or enhanced stability because of slope depressurisation, depending on their location and orientation relative to the pit shape and the water sources. In this study, the shapes of the damage zones and their anticipated impact on infiltration and/or passive depressurisation are considered for different pit morphologies, rock mass conditions, in situ stress, and hydrogeological settings. Some unusual hydrological situations, like mining towards rivers and lakes and the potential effects on pore pressures, are also considered. This investigation has resulted in the development of a method to adjust initial hydraulic conductivity values used to estimate transient pore pressure conditions found in rock masses exposed in different pit morphologies and stress regimes as well as discontinuities’ orientation relative to open faces within the open pit mine. The changes in K values and pore pressures can then be used to determine hydrogeological influence more accurately on future pit slope stability.

Sep 1, 2025

By Piotr M. Kapias

Keywords: extraction, lead refining, Pb-Tl solutions, PbCl2-TlCl solutions. The theoretical part of the paper presents an analysis of the conditions of Tl extraction from Pb using different extractants and determines the activity coefficients in liquid solutions of Pb-Tl and PbCl2-TlCl. The dependence of the solution activity coefficients upon the composition and temperature is presented in the form of Krupkowski and Krupkowski-Fitzner equations. An analysis and evaluation were made of the influence of the methods of PbCl2 introduction into a Pb-Tl solution upon the efficiency of the Tl extraction process under thermodynamic equilibrium conditions. The analysis made it possible to formulate relevant equations describing the dynamics of the process of Tl extraction from Pb under stirring conditions as well as under the extractant injection into the liquid Pb-Tl solution conditions. In the part concerning experimental work, results are given for tests on the Tl extraction process from Pb-Tl liquid solutions of initial Tl content of approximately 0.05% mass using PbCl2 or, alternatively, Cl2(g) conducted in laboratorial scale. The tests made it possible to establish the influence of time, temperature, mass and the method of introducing the extractants into the liquid upon the level of Tl extraction from a Pb-Tl solution. Labels used in the text. WT1m - TI content in metallic phase, % mass WT1s - TI content in chloride phase, % mass wpTlm - TI content in metallic phase in initial state, % mass W°Tlm - TI content in metallic phase in equilibrium conditions, % mass m°m - initial mass in metallic phase mm - mass in metallic phase mPbC12 - lead chloride mass fed to the system MPbC12-TIC1 - mass in chloride phase MPb - lead atomic mass, g/mole MT1 - thallium atomic mass, g/mole K - Nernst distribution coefficient Xi - mole friction of solution component X°PbC12 - mole friction of PbC12 in saturated solution X°T1C1 - mole friction of T1C1 in saturated solution T - temperature, K R - gas constant, J/(mole K) k - kinetic equation constant, 1/(%-min.) ?i - solution component activity coefficient e - extraction rate, %

Jan 1, 2008

By B. Maybee, J. Venter, J. Wesseloo

The selection of open-pit slope angles involves high-value decisions with small changes in slope angle, often representing significant changes in net present value. The existing methods used to define design acceptance criteria have evolved from factor of safety to probability of failure (P[F]) to risk consequence to risk frontiers, which represents current state of the art. Current design acceptance measures are based on references to published tables representing the experience and judgement of their authors, who did not specify the purpose of the measures contained therein. As the purpose of the published design acceptance criteria is not specified (i.e., definite stability, marginal stability, optimization, etc.), such tables cannot be used to achieve optimum slope angles that maximize profitability. This paper develops a design acceptance criterion that maximizes profitability by defining a formulation for optimum risk that balances expected risk and reward. The model developed is titled the Mining Risk Model (MRM), and is applied to open-pit slope angle selection through an equation for optimum probability of downside (P[D]O) that balances the upside or opportunity, the P[F], and the downside impact. This formulation for optimum risk is unique, as many authors have presented objective functions for their risk models that can be optimized, but none of the sources reviewed contained a formulation for optimum risk. The MRM is sufficiently flexible to allow the design performance measure that drives P[F], and hence P[D]O, to be selected based on the intended goal. Furthermore, the essential information that must be known to quantify optimum risk is defined. This allows users to determine what information to collect for optimum risk decisions. A further benefit of the MRM is that slope angle decisions and pit shells can be ranked to select the best option, and a threshold is provided that separates acceptable from unacceptable decisions. Finally, the workflow and information required to determine optimum risk are presented.

Sep 9, 2024

By J. F. Morrison, C. H. Wyndham, N. B. Strydom

INTRODUCTION The shovelling of rock in stopes is still a basic and important operation in many gold mines in the removal of gold-bearing rock from the stopes. There is, however, little factual knowledge of the factors which influence the the rates of energy expenditure and work output, and of the way in which variations in these factors, individually or in combination, affect the mechanical efficiency of the mine workers. What for example is the influence of different rates of shovelling, shovel loads and distances of throw? How does stope width and dip of the stope affect the mechanical efficiency of men shovelling rock? Recently a simple physiological test was introduced into the gold mining industry for classifying Bantu recruits into those capable of hard, moderate and light work. If this test is used to select only those Bantu mine workers who are capable of hard work for the task of shovelling rock in stopes then the rate of energy expenditure which could be expected from these men is 1.5 litres/min for the entire period of the shift. With an adequate knowledge of the influence of the various factors which affect the rate of energy expenditure of men shovelling rock in stopes it would be possible to set work standards for Bantu mine workers in stopes. This should be done in terms of the rate of shovelling, the shovel load and the distance of throw for any combination of width of stope, dip of stope, condition of foot wall and grade of rock in such a way that the resultant energy expenditure would be the 'optimum' that could be expected from the men but would be within their capacities for physical work of an endurance nature (that is, 1. 5 litres/min from the 'average' Bantu mine worker). Work standards could thereby be determined for the different stopes of a mine on a sound and systematic basis. In order to supply the information required a study was carried out by the Human Sciences Laboratory in which the work output and energy expenditures (in terms of oxygen consumption in litres/min) were measured on six highly trained Bantu mine workers under the following conditions: (a) rates of shovelling of 3, 4, 5, 6, and 10 shovels per minute. (b) loads of from 9 to 18 lb on the shovel. (c) distances of throw from 8 to 20 ft. (d) stope width of 28 in., 42 in. and 72 in. (e) dip of stopes of 0°, 15° and 30°. (f) rough and smooth footwalls and various grades of rock size. This paper gives the main results of practical importance of this study. Method The experiments were designed and conducted in conjunction with the Work Study Department of one certain underground conditions such as variable stope widths, stope angle, etc., were made available in a mine Surface Training School and suitable subjects were selected from the large number of labourers being put through the training school. Subjects Bantu recruits who had returned to the mine for a new contract and who had previous underground experience on shovelling were considered to be suitable as subjects for the experiments. The particulars of the subjects are given in Table I. TABLE I PHYSICAL CHARACTERISTICS OF SUBJECTS Weight Height Surface area Max O2 capacity Subject (Kg) (cm) (sq meters) (Litre/minute) A 71.5 167.3 1.80 3.5 B 61.5 162.3 1.65 2.8 C 63.0 164.1 1.69 2.4 D 66.5 167.4 1.75 2.8 E 70.2 164.6 1.77 3.0 F 77.8 173.1 1.92 3.8 Training The method of shovelling, accepted by the Work Study Department, was demonstrated to the subjects who then practised the movements with empty shovels, under the supervision of an instructor. After the technique was mastered, the men started shovelling rock and were trained for at least four hours per day to work at the higher shovelling rates. Training was continued for three weeks to enable the subjects to achieve a reasonable degree of physical training and conditioning. Shovelling method In both the kneeling and standing positions the subjects faced the pile of rock, grasped the shovel in the accepted manner and pushed the blade of the shovel into the base of the pile of rock, sliding the blade between the rock and the concrete floor. In the standing position (72 in. wide stope) the back leg assisted in the movement by pushing the knee against the back of the hand holding the shovel handle. In the kneeling position the elbow was tucked into the side of the body so that the muscles of

By Nelson Sánchez

The Los Bronces Development Project (LBDP) was approved in October 2007 by the Anglo American plc Board at a capital cost of some US$1.74B, and is scheduled to be completed by the end of 2010. The project represents a big step forward for the Los Bronces mine since it will increase the treatment rate by 140%, the mining rate by 64% and almost double copper production during the first four years of the project. This economy of scale will also position Los Bronces as a much bigger and efficient mining operation, better prepared to face adverse market scenarios and allowing Anglo to consolidate its position in base metals. The project addressed numerous scope alternatives such as plant capacity vs mine capacity, cut-off grade strategy to maximise value based on the existing flotation and leaching process, cutback strategy, etc. Main infrastructure aspects of the project are as follows: ?Dump area with 2400Mt capacity for waste and low grade material ?Mining equipment: 1 x electric shovel, 12 x 300t trucks, 1 x drill ?Primary crusher: 60? x 89? ?Three new power lines (144MW capacity) ?Grinding plant at 87kt/day nominal capacity (includes a 4.3km tunnel with a regenerative conveyor belt) ?Grinding plant: 1 SAG mill 40? x 25?, 2 ball mills ?Slurry pipeline (52 km from grinding to flotation plant) with five choke stations ?Flotation plant (rougher, cleaning, regrinding circuits and molybdenum flotation). The history of Los Bronces started in 1867 when it was discovered. In 1978 ENAMI (Government company) sold Los Bronces to Exxon for US$112M. The same year a snow avalanche destroyed San Francisco Flotation Plant. In 1979 the plant was rebuilt at 4.9kt/day and from 1979 to 2002 several plant expansions were made up to 54kt/day capacity. In 2002 Anglo acquired Los Bronces for US$1.3B. The 2002 resource and reserve statement showed reserves at 459Mt at 1.02%CuT and resources at 780Mt at 0.80%CuT. The 2007 statement shows reserves at 1,480Mt at 0.67%CuT and resources at 1,644Mt at 0.45%CuT. Defining the size of a plant expansion is the key challenge that any expansionary project faces. The scope of this paper is a general review of the project, describing all the mining techniques that were used to define the best options that maximise value. In this context, a range of plant capacities (from 58 to 230kt/day) combined with a range of mining capacities (from 83 to 166Mt/a) were evaluated in detail, defining for each option a mine plan, Capex and Opex, and cut-off grade strategy for flotation and leaching processes and the associated Net Present Value (NPV) that was used as the selection criteria. The Los Bronces mineral resources potential provides good growth opportunities for the mine beyond LBDP implementation, offering an exciting invitation to continue studying new initiatives that maximise the value of this deposit.

Jan 1, 2008

By A. W. S. Schumann

A. W. S. Schumann The first reason for todays success lies in the very high standard which was set by our speakers today and I think that we owe a very special vote of thanks to Messrs P. M. C. Wilson, A. A. Oakes, M. J. Deats, A. C. Coetzee, R. E. Burnton and J. G. Ferguson. This, however, cannot account for the fact that there was a tremendous response even before anyone knew who was going to speak or what they were going to speak about. I suggest that the response was enthusiastic because people expected a more informal gathering with a greater opportunity for feedback, for an exchange of views and for questions and answers. The very popularity of the idea has defeated its own ends to some extent. This gathering is too big and many people must have been left with questions unanswered. Possibly one should arrange for a number of meeting places-there are many in the Chamber and the surrounding mining houses-and then have separate meetings in the morning with joint reporting session in the afternoon. Alternatively, we could have a symposium after the colloquium, using Webster's definition of a symposium as as 'a drinking party of merry fellows'. We all know that it is easy to discuss mining into the small hours. Obviously one cannot summarise todays papers and discussions, but I would like to make a few comments on technical matters and after that I would like to speak for a moment about what I see as major trends or tendencies which arise from discussions of this nature. Regarding the question of placing a village on a mine or away from a mine, I would mention that there is quite a body of literature on the adverse sociological effects on the development of children in any community in which the great majority of parents share the same occupation. It is something to be avoided if possible. We discussed the concept of the moving stockpile. If this is proved to be successful, it could save a considerable amount of money for Escom who could then avoid the construction of very large and costly staithes which are now regarded as parts of power stations. One does not see why both a moving stockpile and a staith should be necessary. By the way, speaking of Escom, I see that Mr Bob Scott is here; he has stopped pulling my leg by saying that we sell stone to Escom; perhaps someone has told him that if this is true, it must be the cheapest stone you can buy in this country. My next point relates to the use of epidiascope. It struck me again that this instrument which provides a visual image to accompany your message, can be very effective if it is properly used. These papers speak of a great deal of hard work which went into their preparation, and one really feels that the authors should have taken a few minutes more to prepare their drawings specifically for the epidiascope. There should be nothing on the drawing which is not necessary to illustrate a particular point and there should be nothing on the drawing which is not visible from the back of the hall. In particular, if one cannot read the figures along the sides, then one

By H. C. Voges

An account is given of the tests being conducted by Iscor on the following processes for the beneficiation of coal fines: froth flotation, gravity separation, heavy-medium separation, and pneumatic and column flotation. The coals used in the tests came from the Durnacol, Tshikondeni, Grootegeluk, and Hlobane collieries. The problems involved in the dewatering of fine products are stressed. From a theoretical study of the efficiencies that should be possible for a hypothetical South African coal, it is concluded that future fines-beneficiation plants in South Africa may use standard heavy-medium cyclones for coal down to 1 mm in size, fines heavy-medium separation for the 1 mm to 450 µm fraction, and column or pneumatic flotation for the minus 450 µm fines.

Jan 1, 1991

By K. A. Annan

The last seven hot rolling passes of the ferritic stainless steel AISI430 were simulated on a Gleeble-1500D® thermo-mechanical simulator to investigate the effect of temperature, strain rate, and interpass time on the texture. The compression tests were carried out over a wide range of strain rates (0.1 s-1 to 5 s-1) and temperatures (1100°C to 820°C) with different interpass times (2, 10, 20, and 30 seconds). The dynamic recrystallization to dynamic recovery transition temperature was determined from examining the relationship between the mean flow stress and the inverse of the deformation temperature in multi-pass tests. It was found that the texture in the central layer of the compressed sample is a strong recrystallization-type. The through-thickness textural and microstructural banding, to a large extent, were found to be weakened by hot deformation in the two-phase (a+ ?) region, as opposed to the single-phase (a) region.

Jan 1, 2013

By C. Chapman, M. G. Reading, M. P. Amsden

Computer control of flotation at the Ecstall concentrator By M. P. AMSDEN, C. CHAPMAN, and M. G. READING Ecstall Mining Limited, a subsidiary of Texas Gulf Sulphur Company, operates the Kidd Creek Mine, a 10000 tons per day base metal producer, near Timmins, Ontario. Two different types of ore are mined and milled: copper-zinc and silver-lead-zinc. At the highly automated Ecstall concentrator the ores are processed by selective flotation in three individual circuits. One of the most important metallurgical tools for process control is a sample system incorporating continuous on-stream X-ray analysis. Recently this system was automated further by the addition of computer control of reagents for two copper-zinc flotation circuits. The computer programs, utilizing optimum seeking methods and smelter contract terms, have increased copper recovery by 1,1 per cent, zinc recovery by 1,4 per cent, and have reduced reagent costs by 16 per cent. Although not expected, copper and zinc concentrate grades were improved slightly as well.

By M. B. Cortie

The historical development of materials engineering in the mining industry Is briefly reviewed, and some of the factors currently influencing the industry in South Africa are discussed. Material usages in various parts of the mining operation are surveyed. The predominant mechanisms in the degradation of materials in mines are wear, corrosion, and fatigue. The issue of corrosion-abrasion Is considered, but it Is concluded that abrasive wear predominates in most rock-handling and comminution applications and that, in general, little economic benefit Is to be gained in those applications by the use of existing corrosion-resistant alloys. However, alloys with a measure of corrosion resistance find profitable application elsewhere in the mining environment. The use of other engineering materials such as polymers, ceramics, al1l1inium, and various ferrous alloys may be economically viable in a variety of niche situations. Identification of the most cost-effective material in each case calls for a systematic approach and some patience.

Jan 1, 1996