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

Discussion M. J. Martinson* (Associate Member): This is another interesting paper in the series which the Human Sciences Laboratory is publishing on the physiological cost of the more strenuous manual tasks performed underground. As a mining engineer-albeit presently an academic one -this contributor is more concerned with the practical significance of the study in the field of stope cleaning than with physiological niceties, but before discussing some of the practical issues raised by the paper perhaps the authors would care to reply to lay comment on three physiological matters. Firstly, in two places the authors make statements to the effect that the maximum oxygen intake of 'average' (sic) Bantu mine workers is 3.0 l/min. What do the authors mean by 'average'? Maximum oxygen intake in the individual is probably some function of inter alia age, physical fitness, state of health, ambient partial pressure of oxygen, diurnal and/or seasonal changes in metabolism, and possibly cumulative exposure to dust and other occupational contaminants. Since maximum oxygen intake is apparently such an important quality in manual labourers we ought to have a clear picture of its distribution in the entire labour force, and also know precisely how the individual's quantum is affected by factors such as those mentioned above. Secondly, arising out of the same statements and ignoring the reservation implicit in the previous paragraph, is it correct to say that the 'average' Bantu mine worker has a maximum intake of 3.0 l/min? In the August 1966 issue of the South African Journal of Science four of the present co-authors reported the mean maximum oxygen intake of a random sample of 338 Bantu recruits of different tribes to be 2.63 l/min with a standard deviation of 0.542 l/min, and these figures were repeated in a paper published in the November 1968 issue of this Journal. Admittedly the mean maximum oxygen intake of the six subjects used in the present study was 3.05 l/min, but since they were selected on the grounds of previous shovelling experience they can hardly be regarded as being 'average' recruits. Thirdly, it seems to me that the authors are perhaps a little uncritical in their acceptance of the '50 per cent of maximum oxygen intake' level of activity for endurance work. I am not unmindful of the evidence upon which this level is based, but I suggest that the evidence is at best tenuous and that the whole topic of fatigue and 'optimum' levels of productivity under conditions of high muscular activity and considerable psychological stress is a lot more complicated than would appear from the authors' ready acceptance of the 50 per cent level. Under present working conditions the question is largely rhetorical because it would need an army of supervisors to ensure that all members of a typical gang work at 'optimum' levels throughout the shift, but if individual performance should ever become more readily determinable the issue might become crucial to management and malayisha alike. So much for physiological matters; now a few random thoughts on some wider, practical issues raised by the paper. In the paragraph headed Material the statement is made that 'The mixture (of fine, medium-sized and bigger rocks) was considered to be representative of an average underground rock sample'. Elsewhere in the report the authors specifically draw attention to the effect of particle size on the oxygen consumption of men shovelling graded material, and intuitively one imagines that the size distribution of the material being shovelled would be highly significant in a study such as this. Presumably the authors were forced to assess size distribution subjectively because of the lack of objective data on the subject, and possibly the necessity for this assumption highlights how little we know quantitatively about any of the variables in stope cleaning systems. In this connection it may be remarked that although the present study was largely designed to show how dip and stoping width effect energy expenditure in shovelling, no data are given-nor, probably, are they available-to show how the industry's stoping tonnage is distributed in terms of these two factors. It seems to me that if the mining industry seriously hopes to improve stope productivity it must start by measuring the basic variables in existing systems. Elaborating somewhat on the previous paragraph, there also seems to be a tremendous gap between the present study-and its predecessors-and the prospect of gainfully using the information so carefully garnered by the authors. It would be reassuring to hear from the authors that their work on the physiological aspects is being integrated in some master plan for increasing productivity of manual labour. This I fear is not the case, and if productivity has changed at all in recent years I would hazard a guess that the change has been brought about by random factors rather than by systematic engineering design. The Human Sciences Laboratory has worked extensively in stopes in the past, and possibly the time has come when it should widen the ambit of its operations to include engineering-type analyses of cleaning systems, with particular emphasis on the economics of rapid cleaning by manual labour. Perhaps experiments could be conducted with stakhanovite-type gangs consisting of men selected for their high maximum oxygen intake and superior motivation; these men might be offered special incentives in the shape of extra pay

Jan 10, 1969

By K. Forsberg, Ş. Kaya, C. Dittrich, M. Svärd, E. M. Peters

Scandium can be recovered from waste streams of other metallurgical processes such as red mud from the Bayer process and titanium dioxide acid waste. This is accomplished by an integrated process comprising leaching using acidic media, solvent extraction using organic solvents and stripping of Sc from the organic phase using NH4F solutions. The precipitation of scandium as (NH4)3ScF6 from NH4F strip liquors has been demonstrated at lab-scale with recoveries and purities greater than 98% and 98.3%, respectively. It is desired to assess the possibility of recovering the solvents for re-use in the process to improve its economy and environmental friendliness. Investigations have been conducted to recover alcohol by simple distillation from strip liquor-alcohol mixtures containing about 50 v/v% alcohol. The mixtures were obtained after anti-solvent precipitation of (NH4)3ScF6 using an alcohol to strip liquor volumetric ratio of 1:1. The starting strip liquor contained about 2000 mg/L scandium in 3 mol/L NH4F solution. In the recovery of methanol, the first distillate product collected had a purity of about 87 v/v% and this decreased slightly with time over two hours. The methanol distillate collected in the first 30 minutes with a purity of ca. 85 v/v% was re-used in anti-solvent crystallisation of (NH4)3ScF6 from the initial strip liquor, and the scandium recovery attained was about 98.6 w/w%, almost similar to that of 99.2 w/w% obtained using 100 v/v% methanol. For ethanol recovery, the first distillate product collected was approx. 84 v/v% pure and the distillate collected in the first 30 minutes with a purity of about 75 v/v% achieved a scandium recovery of about 97.2 w/w% in comparison to 98.5 w/w% obtained using 100 v/v% ethanol during anti-solvent crystallisation. Keywords: Alcohol recovery, methanol, ethanol, anti-solvent crystallisation, (NH4)3ScF6

Jan 1, 2020

By S. Gqebe, A. Lewis, M. Rodriguez-Pascual

"Gravitational sedimentation of suspensions in various precipitation processes is hindered by colloidal stability. This is due to the high surface charge of the suspension and results in strong attraction/interaction between the ions on the particle surface and counter-ions in solution. Moreover, this strong interaction results in a charge build-up that renders the suspension stable. In order to induce gravitational sedimentation of these particles, a redistribution of ions close to the particle surface is required. We therefore sought to redistribute ions close to the particle surface by applying a magnetic field. This results in the reduction of interparticle electrostatic repulsive forces and a subsequent increase in the zeta potential of a suspension. For the purpose of this study, a copper sulphide suspension was used. Copper sulphide particles were exposed to a range of field strengths for set exposure times and their zeta potential was measured before and after exposure. All particles had an initial zeta potential value equal to or less than –40 mV prior to magnetic field exposure. A significant increase in zeta potential was observed, with values reaching a maximum of –16.5 mV when exposed to a 2 T field strength for 40 minutes. This is due to Lorentz ion shifts resulting from the Lorentz force exerted by the magnetic field on the particle surface. IntroductionMetal sulphide precipitation is a process for treating industrial waste streams with high residual concentration of mixed metal ions and acid mine drainage. This process has been reported to achieve higher metal removal compared with the industrially prevalent lime precipitation, due to the sparingly soluble nature of the metal sulphide precipitates formed. However, precipitation is governed by high levels of supersaturation which favour the formation of fine particles (Lewis, 2010). Consequently, some of the metal sulphide fines formed exhibit slow settling dynamics and thus tend to remain suspended in solution."

Jan 1, 2016

By M. R. Khalesi, S. Arfania

"Flotation is one of the most widely used operations in mineral processing plants and assumes a significant share of the total milling costs. The purpose of this paper is to introduce a new set of capital and operating cost models for major flotation machines based on the application of single (SRA) and multiple regression analysis (MRA). Thirty-seven major flotation machines were analysed for this purpose. Depending on the machinery type, different technical variables such as diameter, required air flow rate, required floor space, cell volume, required air pressure, and power were considered as predictor variables, individually (in SRA) or simultaneously (in MRA). Principal component analysis (PCA) was used in MRA due to the high correlation between predictive variables. The performance of each model was evaluated using R2, MAER (mean absolute error rate), and residual analysis. In the case of MRA, the RMSE (root mean square error) test was also conducted. Maximum obtained MAER of 13.5% and minimum R2 of 0.86 indicated that these models could be applied as credible tools in estimation of capital and operating costs of flotation machines for design and feasibility studies.IntroductionMineral processing is a vital part of mining projects and mainly involves comminution, sizing, concentration, extractive metallurgical processes, and dewatering. Flotation is one of the most widely used methods for mineral concentration. Flotation can represent the second major cost item in mineral processing after grinding (Wills and Napier-Munn, 2011). Accordingly, it is a main concern of mining project managers to select and optimize flotation circuits in order to decrease costs and increase productivity. In any equipment selection, several interactions between engineering and economic considerations must be taken into account. Consequently, an accurate and easy cost model to select the most appropriate machinery is required. Moreover, cost models could be used in flow sheet simulations applied in design and optimization. Models of unit operations built into the simulators could be improved by linking the equipment cost models (Khalesi et al., 2015)."

Jan 1, 2017

The development of 3D multi-parametric models facilitate the provision of resource information well in advance of the mining face. Using the model, one can evaluate mining areas not only for grade and tonnage predictions but also for predictions of rock mass quality. Blast design and milling requirements can be derived from the rock mass quality predictions. This information can be used for overall mine planning and evaluation, costing, production optimization and slope design. This allows the full range of mining activities and costs to be interconnected, there by lowering costs and improving efficiencies through the application of the multi-parametric model. Feasibility studies for new projects require an expected budget expenditure estimate to within5% of actual costs. This level of accuracy is expected by mining companies and investors but cannot be achieved without a sound knowledge of the rock mass variability. By developing a 3Dmulti-parametric model as part of the feasibility study and using the same geological model exploration boreholes, a significantly better understanding of the rock mass conditions can be obtained. This rock mass information can be used for equipment selection, economic pit layouts, processing plant design, underground mining layouts and support requirements, to name but a few design parameters that depend on rock mass conditions. The initiative at Anglo Platinums, Potgietersrust Platinum Ltd.?s (PPRust) operation, has delivered significant results showing improvements across the mine to mill value chain. These include increased loading rates, reduced electricity and crushing consumables, as well as higher plant throughput. The ore reserve model has gained international acceptance as an invaluable tool for a mining operation. Certainly most financial organizations will not invest in a mining project that does not have an ore reserve model. There is the potential for multi-parametric models to be recognized as vital tools in the mining process, as is the case with ore reserve models.

Jan 1, 2006

Solvent loading: It is agreed that it is advantageous, as far as reagent cost is concerned, to operate at maximum solvent loading. However, owing to the fact that no reliable automatic solvent flow controller is yet available, control tends to be erratic with the result that there is risk of losing uranium. As can be seen from Mr Tunley's plot, the graph representing total ammonia cost is fairly flat between loadings of 3.5 and 4.0 gpl, while the graph representing uranium loss in the raffinate is steep. It is, therefore, safer to operate slightly below optimum loading in the absence of absolute control. Should the purity of the final product gain importance, operation at optimum loading or even slightly above, may be advantageous. Solvent losses and clarification of pregnant solution: The cost of replacing solvent, as indicated by Mr Carman, represents a high percentage of the overall Purlex costs. It is, therefore, agreed that, amongst other factors, efficient clarification of the pregnant solution and mixer-settler design have an important bearing on the cost. However, as at Blyvooruitzicht, the Buffelsfontein plant can tolerate a suspended solids content of 50 ppm without serious deterioration in efficiency. An increase in the suspended solids content from 15 to 50 ppm normally accounts for an increase of approximately 1 ppm in amine loss. Clarification of acidic pregnant solutions is much more difficult than that of alkaline solutions. Sand clarifiers are being used at Buffelsfontein with moderate success. Clarified solutions containing less than 20 ppm of solids in suspension can be produced, depending on the quantity and type of flocculant used in the filtration process, the thickness and the grading analysis of the sand beds and the quality of supervision. An on-stream duty of approximately 0.15 gallon per sq ft of surface area is obtained. Hydraulic handling of clarifier sand, while cleaning, may effect the efficiency of a clarifier, as size segregation of sand particles may lead to inefficient "pockets" in a sand bed. Control of the solution level in the sand clarifier also affects the quality of the effluent. Exposure of the sand bed invariably causes a cloudy effluent. Recycling of the final and initial solution discharged before and after skimming the clarifier sand improves the clarification efficiency significantly. Post-precipitation of silica and possibly sulphate salts, after clarification, may give a false impression of inefficient clarification. This phenomenon is more pronounced when treating hot supersaturated solutions. Ageing and cooling of the solution before clarification assist in stabilising the solution and reduce precipitation after clarification. The maintenance cost of sand clarifiers at Buffelsfontein is in the order of 0.35 cent per metric ton of solution treated. A redundant ion-exchange column has been tested successfully, either as a primary or a secondary clarifier, depending on the quality of the filtrate treated. The conventional pebble and sand beds in the column have been retained. Operation was enhanced by the addition of a two-foot deep resin bed as a filtering medium. The suspended solids content was reduced to less than 5 ppm. Mixer-settlers: Mixer-settler design is a controversial subject but, in view of the high costs of solvent compared to relatively low pumping costs, it is believed that mixing and pumping functions should be separated. Separate facilities for pumping and mixing not only give greater operational flexibility but also permit eventual changes in the flow circuit as technology develops. Scrub technique: Unlike the pilot plant, where the crud was carried away in the aqueous phase, crud is carried over with the organic phase in the full-scale plant. This carry-over of crud into the strip bank aggravates crud formation in the first strip stage. This phenomenon may result from a slight difference in settler design and/or a variation in the relative phase depths in the two plants. Bypassing of the scrub aqueous solution from the fourth to the second stage was introduced in order to avoid dilution and partial neutralisation of the acid added to the third stage, as it is known that a low pH value enhances the transfer of iron into the aqueous phase. The low aqueous flowrate in the third scrubbing stage was overcome by introducing an aqueous phase recycle. Gleaning: Mr MacDonald's contribution on "gleaning" was most informative. This process for recovering solvent from raffinate seems to show at least as much promise as coalescers and other methods employed to date. Similar tests conducted at Buffelsfontein pilot plant gave 60 per cent amine recovery from raffinate containing 10 ppm amine. However, the circuit was not operated at equilibrium and recovery may be expected to drop as the amine concentration of the "Gleaner" circuit increases.

By J. Ning, T. Bu, J-Q. Jiang, L. Jiang

"In coal mining in China, the cutting height of the shearer in longwall faces is increasing. Owing to the increase of extraction height, the caved roof strata area is enlarged, and new issues are being encountered, such as determination of a suitable working resistance for the shield and control of the roof. Through field observation and theoretical analysis of the first longwall face with a height of 6.0 m in the Bayangaole mine, a three-stage structural model was developed. Stage I is defined as the period in which the lower immediate roof (LIR) caves into the goaf and is broken into irregular shapes of various sizes. Stage II is defined as the period in which the upper immediate roof (UIR) breaks and impacts the LIR. In stage III, the main roof breaks into blocks and then impacts the UIR. With respect to these three stages, a suitable method was identified for calculating the working resistance of the shield support for a super-great mining height longwall face (SGMHLW). Introduction and backgroundThe coal seams in the Shendong coalfield, located in the northwest of China, are characterized by low dips and large thicknesses and are thus suited for full thickness extraction. Since the mid-1990s, the Shendong coalfield has been developed into one of the largest in the world (Ning, Liu, and Tan. 2014; Ning et al., 2017). Recent developments in mining equipment have made it possible extract entire seams up to 5.0 m thick in a single cut. This method, termed the ‘super-great mining height longwall face’ (SGMHLW) method, has been widely adopted in the Shendong coalfield (Table I) (Ju and Xu, 2013; Peng, Li, and Zhou 2015; Zhang, Fan, and Ma 2011). According to some studies, sudden support-closure incidents often occur when using the SGMHLW method. The key to solving this engineering issue is to have a clear understanding of the structural characteristics and movement laws of the overlying strata. For instance, the load on the supports is related to the movement of the roof stratum as a response to the mining activities."

Nov 1, 2018

R. T. Whillans and D. A. Cranstone. Canadian reserves of copper, nickel, lead, zinc, molybdenum, silver and gold as of January 1, 1979. Hull (Canada), Canadian Government Publishing Centre, 1980. C$2.40. Brewer, Stephen. Solving problems in analytical chemistry. New York, Wiley, 1980. 512pp. $8.95. E. K. Turkdogan. Physical chemistry for high temperature technology. New York, Academic Press, 1980. 464 pp. $49,50. G. B. Fettweis. World coal resources. Amsterdam, Elsevier, 1979. 416 pp. U.S. $80.50 Dfl. 165.00 Canadian mines: 1979 perspective. Hull (Canada), Canadian Government Publishing Centre, 1980. C$2.40. Clarence Karr (ed.). Analytical methods for coal and coal products, Volume 3. New York, Academic Press, 1979. 656 pp. Charles B. Gill. Nonferrous extractive metallurgy. New York, Wiley, 1980. 500 pp. $30.00. M. Kuzvart and M. B6hmer. Prospecting and exploration for mineral deposits. Amsterdam, Elsevier, 1978. 432 pp. U.S. $68.25. Dfl.140.00. Martin A. Elliott (ed.). Chemistry of coal utilization, supplementary volume 2. New York, Wiley, 1980. 3000 pp. $150.00. M. Radetzki and S. Zorn. Financing mining projects in developing countries. Aedermannsdorf (Switzerland), Trans Tech. Publications, 1979. 200 pp. SFr. 48.00 (hard cover), SFr. 38.00 (soft cover). Charles G. Schofield. Homogenisation/blending systems design and control for minerals processing. Aedermannsdorf (Switzerland), Trans Tech Publications, 1980. 332 pp. SFr.98.00 (hard cover), SFr.48.00 (soft cover). A. V. Smoldyrer and Y. K. Safonov (translator: W. C. Cooley). Pipeline transport of concentrated slurries. Aedermannsdorf (Switzerland), Trans Tech Publications, 1979. 176 pp. SFI'. 115.00. V. M. Karasik (translator: W. C. Cooley). Slurry hydrotransport of minerals and tailings. Aedermannsdorf (Switzerland), Trans Tech Publications, 1979. 156 pp. SFr.1l5.00. S. P. Turchaninov (translator: W. C. Cooley). The life of hydrotransport pipelines. Aedermannsdorf (Switzerland), Trans Tech Publications, 1979. 158 pp. SFr.115.00. NIM reports Report 2037 The removal of contaminants from recycled water in a mineral-flotation plant. Report 2054 The determination, by X-ray spectrometry, of trace amounts of tin in titanium-bearing ores and concentrates and in siliceous materials. Report 2063 Aerosal deposition and carbon-rod atomization of gold. This report describes the determination, by flameless

Jan 1, 1980

The Centenary Distinguished Lecture was held on Thursday, 24th March, 1994 at the Convention Centre of the Sandton Sun, Johannesburg at 20:00. Herewith the proceedings that followed. P.R. Janisch: The South African Institute of Mining and Metalurgy began life one hundred years ago tonight, and eight years (almost to the day) after the 1886 discovery of gold, which has impacted so profoundly on the course of events in our country since then. The Early Years That discovery was a mater of fortune, but not in the sense of the conventional prospector's lucky strike. Canny explorers, such as the Strubens, had been around for some years and, if George Harrison had not stumbled over the Main Reef when he did, someone else would surely have done so within months. The element of luck was that the gold-bearing rocks of the Witwatersrand actually lay exposed to sight. We now know that they are something of the order of 2,5 billion years old. Rocks of that age are usually covered by younger formations. In fact, the bulk of the Witwatersrand Basin is covered by lava, dolomite, and Karoo measures. What, one wonders, would our history have been had the Karoo, which stops at the Vaal, extended no more than 60 kilometres further north. Dr H. Scot-Russell A piece of advice attributed to Paul Kruger, the who was President of the South African Republic one hundred years ago, was the following: 'Seek from the past that which is good, and build on it for the future'. It is advice that can be applied universally and, whether deliberately or not, it must have motivated my predecessors when this Institute celebrated its 40th, 50th, and 75th birthdays. Those milestones were each marked by retrospectives, as Peter Janisch has so ably presented tonight on our hundredth anniversary. The parallels do not end there. We find ourselves in times no less daunting than those our predecessors. Next month will see the birth of a new kind of South Africa; so did May 1902, May 1919, and May 1948.

Jan 1, 1994

By J. K. E. Douglas

PRESIDENTIAL ADDRESS Given by Mr J. K. E. DOUGLAS, M.Se. (Eng.) (Rand) SYNOPSIS Lime is the most widely used and the cheapest chemical alkali known to man and virtually every product we use or eat has required lime in some phase of its manufacture. Its use goes back to the earliest days of man and many ancient buildings and writings bear testimony to this. The manufacture of lime involves quarrying, crushing and screening of limestone and the burning of the sized stone in kilns of which there are several types. The earliest kilns were of very crude design and only in comparatively recent times have large capacity, automated and scientifically controlled kilns been developed. The history of the South African lime industry is largely that of the three main lime companies and their story is briefly told. No other material used in industry has a greater diversity of uses or more varied functions and lime has applications in most South African industries. The main applications are in the production of gold and uranium, iron, steel and ferrochrome, carbide, sugar and paper and for water treatment, agriculture and building. The Republic is well endowed with high quality limestone and the conclusion is drawn that many other minerals will have been exhausted before we run out of limestone with which to process them. The expansion of the lime industry has in the past kept pace with the requirements of industry and the future demand is expected to grow in parallel with the growth of these industries. Since the war this growth has been phenomenal and with our expanding populations and abundance of raw materials it should continue in the years ahead. Limiting factors are the shortage of skilled labour and the distance from export markets. Confidence is expressed that the challenges of the future will be met and that the lime industry will continue to make an important contribution to the growth and prosperity of South Africa. INTRODUCTION Lime, gentlemen, is known to all of you. Like myself, most of you have undoubtedly found it to be, on the appropriate occasion, a most pleasant flavouring for gin. But this lime, or rather its juice, was in earlier days more than just a flavouring. In the days of sail it was vital to the health of sailors as a deterrant against the scourge of all sea-going men-scurvy. It thus played a small but important role in the opening-up of trade routes around the world on which was based the original wealth of many of today's more advanced nations. I propose to address you on the subject of lime this evening, not the citrus variety I have just mentioned, but the most widely-used and cheapest alkali known to man, which, largely unsung, plays an even more vital part in modern industry than did its namesake in the development of inter-continental trade. In the time at my disposal, I propose to tell you something of the history, geology and technology of lime manufacture. Then I will briefly sketch for you the development of this industry in South Africa and indicate the role it plays in supplying the requirements of many of our important industries. Finally we will attempt to look into the future. The importance of this chemical alkali to our modern civilization is seldom fully appreciated-virtually every product we use or eat has required lime in some phase of its manufacture, either directly or indirectly. It is one of those basic materials, along with iron, coal, sulphur, salt and petroleum without which industry and, in fact, our modern way of living could not exist. Fortunately, nature has endowed the world with a plentiful supply of the limestone from which lime is derived and most countries have deposits adequate for their requirements. As a result, not only has there been little trade in lime products between countries but in the past the widespread occurrences of limestone have set a pattern of numerous small plants each serving a limited area. These plants were generally crude affairs incapable of producing a quality product. It is not surprising that, in an industry comprised of small producers with limited financial resources and in fierce competition with each other, there was little interchange of information. Technological development was therefore slow and the industry tended to be regarded as a backyard operation. Its popularity was not enhanced by the fact that lime is difficult and unpleasant to handle if proper facilities are not provided. In recent years the lime industry has undergone very radical changes to meet the more exacting requirements of today's more sophisticated consumers. Only since the war have the larger companies emerged with proper management, highly mechanized plants and a scientific approach. HISTORY Lime was one of the first chemical reagents used by man and consequently lime burning is one of the oldest of the chemical industries. The use of limestone dates back to the stone age when primitive man used limestone to build fireplaces, construct shelters and make crude tools and weapons. It is probable that lime was first discovered by him when the stone in these fireplaces disintegrated to a white powder which he could use for decorative purposes. The first recorded use of limestone was when huge blocks of limestone were used to build the pyramids of Egypt in 4,000 to 2,000 RC. It was not long after this that the beauty of marble, which is in fact a limestone, came to be appreciated and it found wide application in sculpturing and for decorative wall construction. Originally lime, mixed with sand, was used mainly as a mortar for building purposes, the earliest record of its use for chemical purposes being in 350 B.C. in a report of the wreck near Marseilles of a ship carrying a cargo of linen and lime 'for its bleaching'1. Cato mentioned the burning of lime kilns in 184 B.C. and the Romans made extensive use of both limestone and lime for highway construction. A treatise on architecture by Vitruvius who had an official position in the rebuilding of Rome under the Emperor Augustus remained for nearly two thousand

By R. Gebrekristos

It is commonly accepted that three types of aquifers exist in the Bushveld Complex: (a) alluvial aquifers in areas of river courses, (b) weathered bedrock aquifer formed by the in-situ weathering of the bedrocks, (c) fractured aquifers underlying the weathered aquifers controlled by the regional and local fracturing of the crystalline rocks. Hydrogeological studies conducted in the Critical Zone of the Bushveld Complex in the western limb show that the UG2 pyroxenite layer is of a particular interest and has a unique aquifer property. It yields much of the groundwater inflow into some mines and can be classified as the most important aquifer in the vicinity of the mining zone. The UG2 pyroxenite aquifer is restricted to the dipping UG2 pyroxenite unit. The unit weathers more intensely than the surrounding norites and anorthosites and is characterized by deeper weathering horizons (two to three times deeper than the surrounding rocks). For much of the UG2 mining area between depths of 35 ? 80 m, stoping of the UG2 takes place approximately 5 ? 7 m below the UG2 pyroxenite aquifer, and usually there is no groundwater inflow into the mine workings. However, unsealed deep exploration boreholes intersecting the UG2 pyroxenite and reaching underground mine workings, or mine roofbolts penetrating the weathered pyroxenite aquifer from underneath, can result in significant groundwater inflows into the underground workings. The reason as to why the UG2 pyroxenite weathers more than the norites and anorthosites has not been fully studied but is suspected to be due to the bedding planes of the unit and mineralogical composition of the particular pyroxenes forming the unit, which result in preferential physical and chemical weathering under atmospheric conditions. The weathered pyroxenes form sand-sized, rounded grains with improved effective porosity and permeability suitable for the storage and passage of groundwater. The orientation of the pyroxenite aquifer can easily be determined as it dips with the Bushveld layered strata. No-flow boundaries can be assumed on the hanging- and footwall of the pyroxenite aquifer and the flow rates can be estimated using 2-dimentional analytical equations. However, this must be confirmed with strong conceptual models since the existence of a fracture network connecting the pyroxenite aquifer with the overlying weathered and/or underling fractured aquifers can complicate the system, lowering the accuracy of analytical models and necessitating the utilization of numerical models.

Jan 1, 2012

By H. Wagner

The discovery of the gold-rich Main Reef Group of reefs on the farm Langlaagte in 1886 marked the opening of the richest and most extensive goldfield in history. Mining of the narrow reefs progressed from the outcrops to more than 3,5 km, and has made South Africa the foremost mining country in the world. Changes in the political and socio-economic climate create a new challenge to the gold-mining industry. The most significant of these is the emergence of Black trade unions, with their demands for higher wages, improved safety, job advancement, and abolition of discriminatory legislation. Industrial conflicts on gold mines have risen significantly in the past ten years, and are likely to continue in the future, although the causes of these conflicts may change. Events largely outside the control of the gold-mining industry, such as the improved dollar price for gold and the depreciation of the value of the rand relative to major currencies, have enabled the industry to meet the demands for higher wages and to narrow the wage gap. However, labour productivities have improved insignificantly and are a cause of major concern for the future development of the industry. Continued pressure in the industrial-relations area, the inflation of working costs, and the extension of mining operations to greater depths and more difficult areas create major challenges to mine managements and mining engineers. Significant improvements in mining technology, environmental control, and safety will have to be made to ensure the long-term survival of the industry. The unavoidable introduction in deep gold mines of mechanization on a larger scale to meet the requirements of higher labour productivities and a better and safer underground environment will have a major impact on the cost and management structure of these mines. Engineering will assume a more important role on deep gold mil1es, and management styles and structures will have to be adapted to accommodate these changes. The supply of suitably qualified engineers, technicians, and artisans will be a crucial factor controlling the rate of technological change. The industry is well-geared through its Group system to meet these challenges, and to benefit from the experience gained in the operation of highly mechanized mines. However, changes take time to implement and there is very little time left.

Jan 1, 1986

By M. F. Dawson, S. B. Davis

SYNOPSIS This paper describes the size reduction of limestone smaller than 1 mm by attrition grinding, using quartz (1 to 3 mm) as the grinding medium in a laboratory rig. The ore and grinding medium were closely sized fractions, and the particle size of the grinding medium was at least three sizes larger than that of the ore. Batch tests showed that the size reduction of the feed followed two first-order rate processes: one for an apparently fast-breaking ore fraction, and a slower one for most of the ore (86 per cent). The production of fines (material smaller than 10 µm) followed a zero-order rate equation. Continuous runs showed that the process could be modelled as a perfectly mixed system provided that the discharge screen did not selectively retain coarser particles of ore. The ratio of medium to ore (both by mass and by volume) was found to have a significant effect, the most energy¬efficient grinding occurring at ratios between 1 and 2. The solids content of the slurry also had a significant effect, the greatest efficiency being obtained at 50 to 65 per cent solids (by mass). Within these energy-efficient ranges of medium-to-ore and solids content, the lower values resulted in maximum feed breakage and the higher values in maximum fines production. The most important factor in the breakage of feed was the difference in the particle size of the grinding medium and the ore; the amount of ore broken out of the feed size for a given energy input was found to increase linearly with the difference between the size of the grinding medium and that of the ore up to 2 mm. The production of fines increased similarly, but only up to a difference of about 1,2 mm, when it levelled off at about 5 kg/kW ? h. The mechanism proposed for attritioning is mainly the abrasion of 1 µm planes of limestone along the crystal axes, with some fracturing along existing lines of weakness. The energy consumption of nearly 200 kW ? h per ton of product smaller than 10 µm appears to be higher than that in other types of milling. It is concluded that the value of attrition milling lies in its selectivity; in this work, only the softer limestone and virtually none of the hard quartz grinding medium was comminuted.

Jan 1, 1989

By I. Beukes, H. Kotze, B. Krause, R. P. Hattingh

Hillendale Mine is operated by Exxaro KZN Sands, and is situated on the north coast of KwaZulu-Natal in South Africa near the towns of Empangeni and Richards Bay, some 200 km north of Durban. During the mining operation the heavy minerals rutile, ilmenite, zircon and leucoxene are extracted for further beneficiation at a processing complex near the town of Empangeni. The deposit contains approximately 25% slimes (defined as the sub-45 micron fraction), and is mined using hydraulic monitoring. Run of mine is gravity fed down launders to pump stations, which deliver the material to a primary wet plant. Slimes is separated from the run of mine prior to the heavy minerals being separated from the sand. Sand tails containing approximately 2-5% slimes is pumped to the mining void. The slimes is thickened to a density of approximately 1.2, and most of the resultant material is pumped to a sub-aerial deposition site (residue dam). The area is characterized by an annual rainfall of approximately 1300 mm, which is relatively high by South African standards. From an agricultural perspective the soils are deep and well weathered. Dryland agriculture has been practised on an extensive scale in the area for the past 70 years. Sugar cane in particular is cultivated on and around the mine area. In terms of environmental approvals, the post-mining land use and objectives are defined as economically viable and sustainable sugar cane cultivation in the mined areas. During mining the soil profile is destroyed. In this paper the establishment of a reconstituted soil profile is discussed, with specific reference to the methodology that is being followed to achieve the homogeneous blending of sand and slimes. Defining the rehabilitation strategy and soil specification Strategically Exxaro KZN Sands divided rehabilitation into four aspects, comprising physical rehabilitation (restoration of soil structure), vegetative rehabilitation (restoration of vegetative cover), ecological rehabilitation (restoration of sustainable ecological communities), and sustainability (sustainable man/environment interface) (Hattingh and Viljoen, 2006). It is believed that the key to successful rehabilitation lies in achieving a substrate (soil) suitable for the establishment of the target vegetation. During the development of the rehabilitation strategy the key factor in the establishment of a suitable soil profile was shown to be the soil water retention of this material. Since dryland agriculture is practised, this aspect becomes critically important in ensuring that the post-mining land use is sustainable (Hattingh et al., in prep.). From a financial perspective, a closure certificate being issued implies that the state takes over financial liability for the rehabilitated land. This has significant implications for the state-the default position will be that all possible latent and residual impacts must be identified and known prior to a closure certificate being issued. From the mine's perspective this precautionary approach by the state implies that post-operational costs cannot be quantified exactly, and our approach is therefore to mitigate the risk of a long post-closure care and maintenance phase by means of addressing all potential latent defects during the operational rehabilitation phase. The financial viability of the soil reconstitution process itself is therefore not measured against other rehabilitation methods, but against the risk of not achieving closure. During a series of technical workshops comprising experts in the various disciplines in the rehabilitation of the soil profile were identified. These aspects include the re-establishment of a functional soil profile, soil physical characteristics, soil chemistry, soil biota, salinization potential, erodibility and soil weathering. The natural soils in the Hillendale area have been classified in terms of the South African soil classification system (MacVicar and co-workers, 1991) and consist of Hutton (orthic topsoils on red apedal subsoils) and Clovelly (orthic top soils on yellow apedal subsoils) soil forms. The topsoils consistently comprise medium to fine grained sand overlying heavier textured sandy clay loam to clayey subsoils. The topsoil structure consists of a single grain structure, whereas subsoils are apedal. These soils are typically deeply weathered and do not show secondary structures. During the mining process a layer of topsoil (300 mm) is stripped off, and the balance of the material down to the footwall is processed. Sand with 2-5% slimes (defined as the sub-45 micron fraction) is returned as backfill in order to re-establish the dune to approximate the previous landform. From an agricultural perspective this material is not suitable for the dryland cultivation of sugar cane. A capping of more suitable material therefore has to be provided. The most significant difference between the premining soil form and the backfill material is the percentage slimes, consequently a mixture of the sand and slimes is required as a capping over the backfilled sand. The required thickness of the cap is a function of specifically plant growth requirements, erodibility, weathering and runoff (Hattingh, et al.; in prep.). The percentage slimes required in the cap is a function of mainly soil water requirements (Hattingh and Viljoen, 2006). The conservative indications to date show that the reconstituted soil profile should consist of a homogeneously blended mixture of 70%-80% sand and 30%-20% slimes up to a maximum of 2 metres thick. This will be sufficient to ensure growth during dry periods.

Jan 1, 2007

By D. Moradkhani, S. M. Javad Koleini, M. Abdollahy, F. Soltani

"The objective of the study was to determine the type and distribution of rare earth elements (REEs) in the phosphate flotation concentrate from the Esfordi flotation plant in central Iran, in order to understand their leaching behaviour. Samples of the concentrate were prepared and analysed using XRF, XRD, ICP, and EMPA. The Ce, Nd, La, and Y contents were 5608, 2227, 1959, and 679.7 mg/kg, respectively. The main REEbearing minerals in the concentrate are fluorapatite, monazite, and xenotime. Apatite is the chief mineral of the concentrate and the ratio of apatite to monazite is greater than 60. From EMPA analysis, monazite contains a much higher proportion of REEs than apatite, but more than 50% of the REE content of the concentrate is in apatite, as this is the main mineral constituent of the concentrate. During the leaching of REEs in the phosphoric acid production process, more than 99% of the total rare earth values report to the large amounts of phosphogypsum (CaSO4·2H2O) formed during the dissolution reaction. Accordingly, two process options are suggested for treating the Esfordi concentrate. In the first option, the concentrate is treated with 98% H2SO4 at 200–300°C. After digestion and water leaching, phosphorus and REEs are precipitated out of solution. In the second option, the concentrate is treated in a pre-leach stage. The process residue from the pre-leach stage is then subjected to acid digestion at elevated temperature for phosphoric acid production.IntroductionThe rare earth elements (REEs) comprise a set of 17 chemical elements in the periodic table, specifically the 15 lanthanides plus scandium and yttrium. REEs and alloys that contain them are used in devices such as computer memory, rechargeable batteries, cell phones, catalytic converters, magnets, fluorescent lighting, and many more (Krishnamurthy and Gupta, 2004). China supplies about 94% of the REE demand, with the remaining 6% coming from Russia and Estonia, the USA, India, Malaysia, and Brazil (Zhanheng, 2011). Increased industrial development in China has prompted the Chinese government to limit annual export quotas to approximately 35 kt of rare earth oxides (REOs), while non-Chinese annual demand is expected to reach 80 kt by the year 2015. This constriction of supply is being met by the development of many new rare earth mining projects, each of which has its own unique mining and processing challenges (Jordens, Cheng, and Waters, 2013). One of these projects is the Esfordi phosphate project, which is located in Yazd Province in central Iran. The concentrate from the Esfordi flotation plant contains 1.2–1.5% REEs."

Jan 1, 2017

By R. Hewer

Transition metals have been widely used in medicine throughout history. The medicinal use of metals extends back to the earliest recorded history when the Chinese used gold as treatment for a variety of ailments. The ancient Egyptians used copper to sterilize water and mercury was reportedly used by Hippocrates in 400BC. Thereafter, numerous other metals have been used as treatment throughout the centuries. For example, zinc sulphate was used as an emetic in the 17th century and zinc oxide was used to treat epilepsy in the 18th century. This use of metals, and numerous other instances, which have been well described in Sneader 2008,1 make for interesting reading but largely occurred prior to the advent of rigorous scientific method. Therefore, although there are exemptions, most therapeutic applications of metals prior to the modern age lack scientific grounding and have been found to be mostly ineffective. Of more relevance are metal-based medicines approved by regulatory bodies (i.e. Food and Drug Administration; FDA) that are in current use as therapeutic drugs. There are presently a large number of metal-based drugs that are approved for use by the FDA ranging from over the counter medicines, such as bismuth subsalicylate for gastrointestinal problems, to prescription drugs such as lithium carbonate for the treatment of manic depression. The list of metal-based medicines approved by the FDA not only incorporates base metals but also includes precious metals. Platinum complexes have been found to be potently active against several different types of cancers while gold complexes have shown application as treatment for rheumatoid arthritis (RA). With the well-established foundation that exists for metal-based medicines, it is imperative to continue exploration into new therapeutic areas. The precious metals; gold, ruthenium and platinum, continue to attract scientific interest in the field of biomedicine and have been widely researched. Scientific studies conducted on gold complexes have demonstrated the potential of this metal for the treatment of HIV/AIDS, cancer and malaria. These diseases are also specific health concerns for developing nations; countries where disease has devastating impact. Within South Africa alone, an estimated 5.7 million people are infected with HIV/AIDS,2 approximately 7 000 new cases of malaria are reported annually, and it is projected that 1 in 4 persons will develop cancer. Continuous development of therapeutic drugs to combat these diseases is therefore apriority, and the exploration of gold-based compounds for this purpose is warranted.

Jan 1, 2009

By M. Tang, X. Tong

"SynopsisSlime coating is one of the most common ways for serpentines to contaminate metallic mineral concentrates during traditional flotation of coarse sulphide particles. This could pose quite a complicated and challenging problem in the case of some types of low-grade and finely disseminated Cu-Ni ores bearing high serpentine contents. This is the case for the copper and nickel sulphides from the Yunnan Mine, China. Previous batch flotation tests of this ore resulted in satisfactory recoveries of 86.92% Cu, 54.92% Ni, and 74.73% Pt+Pd, and concentrate grades of 4.02% Cu, 3.24% Ni, and 76.61 g/t Pt+Pd. However, the MgO content in these concentrates was more than 19%. In the current study, microflotation tests and settling rate tests were introduced to investigate the effects of a combination of strong collectors (a 2:1 weight ratio of butyl xanthate and butyl ammonium dithophosphate) on entrainment of serpentines in metallic mineral concentrate, as well as visual observations of the concentrates in suspension using still photography. All test results indicated the presence of serpentines entrapped in the hydrophobic flocs that resulted from these collectors, even with the use of effective gangue depressants. These strong collectors are used to flocculate the ultra-fine sulphides by forming loose and ‘fluffy’ hydrophobic flocs. However, these hydrophobic flocs may also be able to load or entrap some serpentine slimes into the concentrate, and this entrained serpentine could be harder to remove by using depressants or intensified conditioning than serpentine slime coating on the particle surfaces.IntroductionSerpentine minerals, which have the generalized composition (Mg,Fe)3Si2O5(OH)4, can be easy to crush and grind due to their convoluted and bent layered structures. These serpentines can hinder the enrichment of some metallic minerals and dilute their concentrates by entrainment during flotation. The resulting high pulp viscosities, slime coatings, and high content of dissolved ions adversely affect the flotation recoveries of copper and nickel sulphides, and high levels of MgO entrained in the concentrates lead to heavy penalties from smelters. Since the quality of nickel flotation concentrate depends heavily on its MgO content (generally less than 6–7% for No.1 or 2 grade nickel concentrate according to the standard requirement from National Nonferrous Industry), reducing the entrainment of serpentine during flotation is becoming more and more pressing. Previous research has focused on the collectors and depressants that are commonly used in copper-nickel flotation, as well as their interaction mechanisms."

Jan 1, 2015

By H. W. J. P. Neomagus, J. R. Bunt, T. S. Mthombo, C. A. Strydom

"Low-cost calcium lignosulphonate (CL), which is a byproduct of the wood pulping industry, was investigated as a potential binder during laboratoryscale pelletization of inertinite-rich high-ash coal fines from the Highveld Coalfield of South Africa. Coal-binder mixtures containing between 1 and 15% calcium lignosulphonate were prepared, pressed, and ultimate, proximate, porosity, XRD, and XRF analyses obtained. It was found that increased pressure during pelletization results in increased attractive forces, due to increased interparticle contact areas. FTIR analysis showed that chemical interactions occur between coal fines and the binder, as evidenced by the formation of new C-O ether-type bonds. The presence of moisture during pelletization enhanced surfactant activity of the binder. Chemical interactions between the coal and the binder seem to be mainly electrostatic, driven by Ca2+ from the calcium lignosulphonate, which may act as a bridge between the anionic parts of the binder and the surface of the coal. IntroductionCoal has over the years remained South Africa’s major energy source, with over 90% of the country’s electricity being produced via coal-fired processes (Arndt et al., 2016). Sasol’s coal conversion technology alone provides more than 40% of South Africa’s liquid fuel requirements (van Dyk, Keyser, and Coertzen, 2006). The production of high-grade coal requires the processing of run-of-mine (ROM) resources by a combination of processes which include crushing, screening, and washing (Radloff, Kirsten, and Anderson, 2003). These processes result in the generation of coal fines (< 0.5 mm) which, until recently, have been discarded into slimes dams (le Roux, Campbell, and Smit, 2012). Several issues are associated with slimes dams, among which are the costs of transportation of these fines from the place of origin to the discard point (often a considerable distance), and the potential environmental effects due to seepage of toxic materials into surrounding ecosystems (Clausen, 1973).Consequently, utilization of these coal fines is an area of great interest. Mechanical densification of fine coal by means of pelletization and/or briquetting to larger sizes for use in technologies that require lump coal, such as chain-grate stokers and fixed-bed gasifiers, has been shown not only to reduce environmental risks (Mangena et al., 2004), but also to significantly reduce storage and transportation costs (Mani et al., 2001). Pelletization is the process by which solid particles are consolidated into uniform, usually hard, impact-resistant agglomerates, rendering them useful as a fuel with sufficient calorific value for domestic and industrial utilization (Finney, Sharifi, and Swithenbank, 2009)."

Dec 1, 2018

By C. F. Vos, J. D. Miller, J. C. Davidtz

"This work presents the potential for improving the flotation recovery of slow-floating sulphide minerals with the use of starvation dosages of a normal dodecyl (n-C12) trithiocarbonate (TTC) co-collector, together with a sodium isobutyl xanthate (SiBX) and dithiophosphate (DTP) collector mixture.At potentials below –150 mV (SHE), addition of nC12-TTC with SiBX improves the hydrophobicity of pyrrhotite, yielding captive bubble contact angles greater than those measured for SiBX or nC12-TTC alone, suggesting a low potential synergistic effect. This synergistic effect is further studied using Fourier transform infrared (FTIR) spectroscopy, the results indicating an increase in the surface concentration of the collector species when in a mixture. Thus, nC12-TTC with SiBX may act as an immobile surface anchor to which SiBX/SiBX2 molecules bond, increasing the localized concentration of collector species.Bench-scale flotation tests using mixtures of SiBX/DTP/nC12-TTC on a platinum group element (PGE)-bearing ore from the Bushveld Complex in South Africa confirm an improved metallurgical performance at very low substitutions (approx. 5 molar per cent) of SiBX. The improved recoveries for PGE, Cu, and Ni are correlated with improvements in the flotation kinetics of their slow-floating components. IntroductionShort-chain (less than C6) xanthates are very effective for the bulk flotation of sulphides (Fuerstenau, 1982), and early work (Leja, 1968; Gaudin, 1957; Plaksin and Bessonov, 1957; Taggart, Giudice, and Ziehl, 1934) indicated that the presence of oxygen in flotation slurries is necessary for the oxidation of the xanthate collector at the sulphide mineral surface to induce hydrophobicity. Fundamental research into the chemistry and adsorption mechanisms of xanthates over the years (Finkelstein and Poling, 1977; Woods, 1976; Winter and Woods, 1973) has revealed that for iron-bearing sulphide minerals, adsorption occurs predominantly through a charge transfer process. Oxidation of the adsorbed collector to the corresponding dimer takes place if the mixed potential of the system is greater than the reversible potential for dimer formation. In oxygenated slurries the rate of adsorption and oxidation of the collector depends greatly on the substrate surfaces. For chalcopyrite (Guler et al., 2005; Leppinen, 1990; Roos, Celis, and Sudrassono, 1990) and pentlandite (Hodgson and Agar, 1989) a two-step interaction with xanthate is proposed in which the chemisorbed xanthate is further oxidized to the dimer. The initial interaction of xanthate with pyrrhotite is through physisorption where xanthate physisorbs onto positive surface sites followed by oxidation to dixanthogen (Khan and Kellebek, 2004; Bozkurt, Xu, and Finch., 1998). This process is, however, slow and increased reaction time is needed to improve flotation recovery (Buswell and Nicol, 2002)."

Oct 1, 2018

By J. N. Van der Merwe

Following the Coalbrook disaster in 1960, research into coal pillar strength resulted in the adoption of the concept of a safety factor for the design of stable pillars in South African coal mining. At the time when the original statistical analysis was performed by Salamon and Munro in the early 1960s, 27 cases of failed pillar workings were considered suitable for inclusion in the database of failed pillars. Pillar failure did not stop after the introduction of the safety factor formula by Salamon and Munro (1967). In the ensuing years, pillars that were created before the application of the formula deteriorated and later failed, as did ones that were created after the introduction of the formula. This means that over time, the database of failed pillar cases increased in size, allowing ever more reliable analyses to be performed. The number of failed cases in the database had grown from the original 27 in the 1960s to 86 by 2011. All the failed cases are contained in the updated database. The database of stable pillars, which is also used in the derivation of strength formulae, has now been extended from 125 to 337 cases. The new database of intact pillar cases is more complete as it bridges the time gap between the Salamon and Munro (1967) and the Van der Merwe (2006) databases. The original requirements for inclusion into the database were satisfied in the compilation of this latest collection. The characteristics of the original database of intact pillars did not change in a meaningful way. The mining depth and pillar dimensions of the new database are largely as they were in the original database. Time-related trends with regard to pillar dimensions and depth of mining could not be found, indicating that the geometrical parameters of coal mining in South Africa have not changed meaningfully in approximately a century of mining. The characteristics of cases in the updated database of failed pillars does not differ substantially from the one published by Van der Merwe (2006). The same difference between that database and the original Salamon and Munro database, namely that the average safety factor of the failed cases had increased dramatically, from 1.0 to 1.5, is still apparent. This may be due to the inclusion of more failures from specific areas that exhibit a disproportionate number of failures at higher safety factors. These areas are the Vaal Basin, Klip River, and Free State coalfields. The new database confirms yet again that there is no correlation between the safety factors of failed pillars and their time of failure. The safety factor on its own is thus not a reliable predictor of longterm stability of pillars. Keywords coal pillar, stable pillar cases, pillar collapse, pillar data base.

Nov 1, 2013