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By G. V. R. Landman

In the early days of underground coal mining, the use of explosives to break coal resulted in a large number of disastrous methane and coal dust explosions and consequential loss of life. Research by the British, French, US authorities, amongst others, introduced many practices that significantly reduced the occurrence of such incidences. The use of explosives designed to be safe in a methane atmosphere, plus the creation of a second free face were two such practices which became commonly used in coal mines all over the world including southern Africa. With its associated coal cutting equipment and labour, the cost of creating a second free face is significant and this, together with the constant drive to reduce mining costs in a competitive market, causes mine management to regularly question the need for the second free face. Also, the safety and nature of permitted explosives has changed considerably since the days of black powder and nitroglycerine-based explosives with the introduction of water-based explosives. In southern Africa mining regulations pertaining to the blasting of coal has its origins in British research which was largely done in the first half of the twentieth century. More recent research by the US Bureau of Mine has resulted in a deeper understanding of some of the mechanisms of methane ignition by explosives and suggests that some of the earlier regulations can be relaxed to a certain extent without compromising safety. This paper gives an overview of the original research done in Europe and the more recent work in the USA, to give mine management and regulatory authorities a deeper understanding of the important safety issues when using explosives in a methane atmosphere. It also suggests that blasting off the solid is no less safe than using a second free face providing certain safety procedures are followed.

Jan 1, 2000

By P. J. Bhonde, R. D. Angal, R. H. Tupkary

Manganese imparts numerous beneficial mechanical properties. Hence it finds extensive application in iron, steel and non-ferrous sector. Prevalently, it is incorporated in the form of master alloys i.e. 'ferromanganese'. In India, ferromanganese is mainly the inexpensive and paramount source of availability of manganese. Various conventional and non-conventional techniques of ferromanganese production have been adopted in past and present. However, these techniques have demerits viz. high refractory, energy consumption, losses of valuables in slag and fumes. Also geographical co-ordinates of the plant restricted the manufacturer to prefer an appropriate process. Accordingly, the industrially implemented methods had reliance on availability of infrastructure, energy, mineral resources, grade of raw material, outlay of labor, environmental compatibility of technique etc. In view of the liberalized economy and the gap scenario in the supply and demand of ferromanganese, the authors felt it indispensable to explicitly bridge up this vacuum by initially designing a simple energy conserving, eco-friendly, low capital investment technique for decarburization of ferromanganese. Measurement of the data involved studying the impact of rate governing parameters on the process kinetics. This laboratory scale investigation is expected to escalate the Mn/C quotient of high carbon ferromanganese. Key Words: ferromanganese, carbon, carbide, decarburization, rate, kinetic

Jan 1, 2007

By G. Sahasrabudh, G. DeIulii, K. P. Galvin

A novel agglomeration process has been developed, utilising a concentrated water in oil emulsion as a binder. The binder is typically 95 wt% water, the remainder being an organic phase containing kerosene and the emulsifier— sorbitan mono oleate (SMO) in equal mass portions. The binder can selectively agglomerate hydrophobic particles in less than 7 seconds. The agglomeration has proven to be highly effective in recovering any particle of virtually any size, provided the particle is hydrophobic. The approach has previously been applied to coal, magnetite and silica. In the paper, sodium isobutyl xanthate (SIBX) and its much longer chain version, potassium tetradecyl xanthate (C14X), were used separately as collectors to selectively agglomerate chalcopyrite from the ore. Xanthates attach to metal sulphides, thus hydrophobising sulphide mineral particles. Chalcopyrite is a sulphide mineral; however, it hydrolyses quickly in the presence of water. Ethylenediaminetetraacetic acid (EDTA) was used to remove the oxide layer formed at the surface of the mineral particles after wet grinding, leaving behind a freshly generated metal sulphide surface. After the EDTA treatment, the ore particles were quickly conditioned at an optimised pH as well as collector concentration and agglomerated. The agglomeration achieved a copper recovery of 75% within 7 seconds, consistent with the recovery achieved by flotation over much longer time scales. Keywords: Chalcopyrite, agglomeration, froth flotation, interfacial chemistry, X-ray fluorescence, selective hydrophobisation

Jan 1, 2020

By G. M. Evans, S. Mitra, G. J. Jameson, M. M. Hoque, E. Doroodchi, R. Moreno-Atanasio

In this study, the detachment behaviour of 314 μm diameter silica particles from a single 3 mm diameter bubble rotating in a confined cavity connected to a cross-flowing water channel was investigated. The liquid cross-flow generated a rotating flow inside the confined cavity, which was quantified using 2D particle image velocimetry (PIV) measurements. The fluctuating (x-y) velocity components inside the cavity for each cross-flow velocity, U, were then used to estimate the corresponding time/volume average turbulence intensity values, Ti. Turbulence intensity ranged from 9.3-16.2%. In the absence of a bubble, the rotating vortex within the cavity created a pressure field which decreased towards the centre of the vortex; wherein the lowest pressure zone was located at the centre of the cavity. The vorticity into the cavity was found to increase with increasing the turbulence intensity. A bubble, both without and with particles attached, was introduced into the cavity under different flow conditions and its motion was tracked using the shadowgraphy technique. Experimentally, it was found that the bubble-particle aggregate was trapped at the centre of the cavity (low-pressure region) following a spiral trajectory. An interface resolved 3D computational fluid dynamics (CFD) model was developed to predict the particle detachment conditions within the confined cavity. CFDpredicted vorticity in the absence of a bubble agreed well with experimental observation. The CFD analysis indicated that the vorticity inside the cavity was not significantly changed when a single bubble was introduced. On this basis, the computed vorticity was utilised in describing the force balance applied to the attached particle. Computationally, it was found that the critical turbulence intensity, beyond which the particle was detached from the bubble, was 14%. This value was consistent with the experimental observation of detachment at Ti at 15.7%, and above, but not at lower values. Keywords: Bubble, particle, turbulence intensity, vorticity, particle detachment, CFD, PIV

Jan 1, 2020

By R. P. Mohring

The South African Coal Industry had its beginnings in the Eastern Cape in 1859 when coal was mined to satisfy the fledgling settlements in the Eastern Cape. The growth in the industry was relatively slow up to the early 1970s when, firstly, the export business was expanded coupled with a major increase in the quantity of coal converted to liquid fuels and utilized for power generation. To meet this increased demand the industry developed many new operations employing the latest mining technology and systems. Coal production in South Africa reached 204 million tons in 1996 and revenue from exports reached R8 billion making it the second-largest foreign exchange earner after gold. In his address Rick P. Mohring describes the role coal plays in the international energy scene and outlines South Africa?s position as a producer and consumer of coal, and the challenges facing the industry. A wide range of steam-coals are produced to satisfy both the local market (power generation, coal conversion and the metallurgical industry) and the export steam-coal market. South Africa is in a unique position regarding the export market as it is still a relatively low-cost producer and is geographically well-placed to play in both the European and Far East markets. There are, however, a number of challenges facing the industry. To retain our competitiveness in the international market, ways of minimizing the cost pressures caused by high-cost inflation, increased environmental protection obligations and a decreasing quality resource base need to be sought. Locally the coal mining industry has a major role to play in Eskom?s goal of reducing the real cost of electricity by 15% in the next three years. Safety and health, productivity and the adaptation of the latest technology are issues being addressed by the industry to maintain its dominant position within the South African economy.

Jan 1, 1997

By J. H. French, O. E. Lissner

S. J. Venning (Member): The authors should be congratulated on producing such concise data from the plethora of experimental report results. Their observations on the cost of liners calls for some comment however, as the really unique feature about liners in South Africa is that they can be obtained so cheaply. Attention has been focussed on the 1961 cost of white iron at R50 per ton for chill cast block liners while manganese steel liners are reported as costing as much as R250 per ton. The former are however, a relatively heavy, simple casting with a low labour content and reasonably high percentage productivity, made from metal melted in a cupola. The capital investment to produce this type of casting is low compared with that required to produce manganese or other alloy steels. The chill cast liner block is always a production item and is cast in a permanent mould. Almost all manganese mill liners, except perhaps small aperture screens, are sold for considerably less than the R250 per ton quoted by the authors. Where a price of this nature is asked it would be because the design of the casting demanded is complicated or leaves little metal to carry the labour and other costs of production. As reference has also been made in the paper to liner costs in America, prices for similar types of liners obtained during a visit to Canada and the U.S.A. in 1961, could be of interest: Material Price per Location short ton Ni-hard R250 Elliot Lake Body liners block type Ni-Hard R300-310 FAS New York Mn Steel R280-290 FAS New York Screens (grids) slotted Cr.Mo. Steel R600 FAS New York hole Ni-Hard R320 Elliot Lake At this time the price of comparable manganese steel block type liners on the Witwatersrand was between R145/R155 per short ton. White iron in Canada was then quoted at about R150 per ton, but as Ni-Hard gave a life of at least two to twoand-a-half times that of white iron, it was a much more attractive proposition particularly when relining and down time costs were taken into consideration. There is little doubt that most foundries would welcome clear and complete specifications of the customers requirements and systematic checks to ensure that these specifications are being met. B.S.S. or other accepted standards should form the basis of these, and periodic inspections should be undertaken to ensure that only those foundries capable of complying should be included as suppliers. The authors have referred to new designs, but can the industry really call the grid and plate new when the patents were taken out during the second world war and the sawtooth block has been used for at least 35 years. Unfortunately, far too little thought and consideration are normally given to the ultimate mill liner design while the mills themselves are on the drawing boards. The position and size of the door opening, which is the key to the ultimate shell liner configuration, the drilling of the

Jan 9, 1968

By M. D. Lionnet

The Bafokeng Rasimone Platinum Mine (BRPM) is a joint venture between Rustenburg Platinum Mines Limited and Royal Bafokeng Resources (Pty) Ltd (Bafokeng). The mining lease has been granted over Boschkoppie 104JQ and is located approximately 10 km south of the Pilanesberg Alkaline Complex. The Merensky Reef, which constitutes the principal orebody, is geologically complex. A relationship between its geological variability and the associated mineralization has been established by defining a number of facies types. The Merensky Reef has been subdivided into seven distinct facies types: the pegmatoidal facies, contact facies, FW1 potholefacies, FW3 pothole facies, FW6 pothole facies, and FW7 pothole facies. The Merensky Reef, which does not form part of the potholed reef, has a large component of footwall mineralization. The pothole facies, in contrast, has mineralization located in the hangingwall. Statistical analysis of the facies types indicates that they are unique populations and hence need to be evaluated independently. The mining lease has distinct structural domains that have added additional complications to evaluation of the Merensky Reef. The Merensky facies on BRPM exhibit unique geological and statistical characteristics within well-defined structural blocks. Due to the extreme variability of the Merensky Reef on BRPM a unique approach has been developed, within Anglo Platinum, to evaluate the deposit. Mining cuts are defined and composited for the individual facies types, and then modelled individually by applying known geological controls. The resultant grade models are weighted based on statistical and local geological knowledge. The object of weighting the facies is to replicate the variability observed in the orebody. The introduction of stringent geological controls to the evaluation process has resulted in increased confidence in the in situ grades estimated as an input to the mine planning process.

Jan 1, 2007

By A. V. Roschin, V. E. Roschin

Composite ores are more often used for production of iron in connection with reserve depletion of traditional ore crude in the regions with developed metallurgy. In particular, the majority of perspective ferrum-ore deposits of the Urals are presented by titanferous magnetite ores which contain a significant amount of titanium and vanadium alongside with ferrum. Thus, the deposits of titanferous magnetites of the Southern Urals on the average contain about 20 % of Fe, 8,7 % TiO2 and 0,24 % V2O5. Due to a high content of titanium oxides it is impossible or it is inexpedient to process such ores in blast furnaces. During work just on titanferous magnetite ore refractory titanium-containing slags are formed in a blast furnace, and in case of its dilution by ores of different types the titanium due to the small content in slag is thrown out. It is often impossible to separate such ores into ferriferous and titanferous concentrates because ferrum and titanium are present in the same minerals. This work attempts to selectively reduce ferrum in solid oxide and to separate metallic iron from titanferousslag with fast fusion. Solid-phase reduction of ferrum has been carried out with carbon at temperature of 1000? 1300°?. In products of solid-phase reduction the extremely disperse mix of metal iron and only partially reduced titanium oxides have been achieved. It is impossible to separate such a mix mechanically. However, the fast fusion of the mix allowed separating it into cast iron and slag with the high content of titanium. Thus the cast iron contained less than 0,1 % of titanium, and the slag did not contain practically any ferrum.

Jan 1, 2007

By S. Ranganathan

Production of ferro-chromium in the submerged-arc furnace is highly energy-intensive. Several process rates have been attempted to reduce the specific energy consumption in the production of this alloy. An exhaustive understanding of the thermo-chemical features of the production process can be used to reduce the specific power consumption significantly. The knowledge can be used to reduce the energy drain through undesired reactions in the process can be minimised. Plant trials have shown that specific energy consumption up to 100kWh/tonne of hot metal can be obtained by adopting this strategy. The process requires modification in operation practice and does not require any capital investment unlike other processes adopted for reducing power consumption. The principle of thermo-chemistry involved in this process development and results of plant trials are discussed in detail.

Jan 1, 2007

By A. McGarr

The shear stress t that can be sustained by the rock mass in the environs of a mining-induced earthquake controls the near-fault peak ground velocity v of that event according to v=0.25(ß/G)t, where ß is the shear wave speed and G is the modulus of rigidity. To estimate t at mining depths, I review the results of four studies involving Witwatersrand tremors that relate to the bulk shear strength. The first and most general analysis uses the common assumptions that the seismogenic crust is pervasively faulted, has hydrostatic pore pressure before mining, and an extensional stress state that is close to failure. Mining operations reduce the pore pressure to zero within the mine and redistribute the stresses such that, in localized regions, the state of stress is again at the point of failure. Laboratory friction experiments can be used to estimate t in the zero-pore-pressure regime. Second, model calculations of states of stress in the vicinity of mining at about 3 km depth indicated the shear stress available to cause faulting near the centre of a distribution of induced earthquakes. Third, laboratory experiments combined with microscopic analyses of fault gouge from the rupture zone of a mining-induced event provided an estimate of the average shear stress acting on the fault to cause this earthquake at a depth of 2 km. Fourth, moment tensors determined for mining-induced earthquakes usually show substantial implosive components, from which it is straightforward to estimate t. These four different analyses yield estimates of t that fall in the range 30to 61 MPa which implies that near-fault particle velocities could be as high as about 1.5 m/s. To the extent that the causative fault ruptures previously intact rock, both t and v, in localized regions, could be several times higher than 61 MPa and 1.5 m/s.

Jan 1, 2002

By J. L. Human, N. Fernandes

The rockfall hazard is the major contributing agency to fatal accidents in South African platinum mining operations. The majority of these accidents occur in close proximity to the working stope face. The collapse of fractured or blocky ground between individual temporary support units or the closest permanent support line to the stope face has not yet been adequately addressed. A large-scale stope support system testing facility funded by the Safety in Mines Research Advisory Committee (SIMRAC), has been designed and constructed to test support systems and stimulate innovative thinking by mining personnel, support manufacturers and suppliers by providing a facility where the interaction of stope support systems, a rigid stope face and a discontinuous hangingwall beam can be observed during quasi-static and dynamic conditions. Tests have been carried out on various types of stope support systems, focusing on the temporary support systems used in the immediate stope face area. The interlinking of the temporary support units with mesh or safety netting has been explored in the prevention of falls of ground between support units during gravity driven falls of ground. These tests illustrate the facility?s ability to be used to stimulate innovative thinking and provide a testing ground for the development of temporary support systems in order to reduce the risk off all of ground injury to mining personnel at the stope face.

Jan 1, 2004

By M. I. Brittan

Discussion Dr R. E. Robinson (Fellow): The author must be congratulated on a very meticulous and self-contained piece of work. It is indeed a pleasure to read a paper that is so clearly and systematically laid out, and where the conclusions and the testwork conducted have been so clearly described. The paper is complete in itself, which makes it very difficult for someone who is not intimately involved in the whole Torco project to make any comments on its content. However, two points that, strictly speaking, fall outside the scope of the paper are of considerable interest. The first relates to the particle size of the material treated. In the paper, the testwork is confined to one standard particle size (minus 60 plus 100 mesh). The essential feature of the paper is to indicate that the rate-controlling reaction in the whole segregation process is the rate of reaction of the ore particles with the reducing agent and the hydrogen chloride. One wonders, therefore, to what extent this relatively slow rate of reaction is affected by the particle size of the ore itself. One imagines that the reaction must take place by contact of the hydrogen chloride with the surface of the mineral particles, and it is reasonable to suspect that the rate of diffusion of the copper ions to the surface is a relatively slow process and is thus the limiting factor in this particular rate of reaction. It is possible, for example, that the improvement obtained, when the ore is subjected to reducing conditions before the chlorination, is due to a breakdown in the crystal structure of the original particle. This breakdown is brought about by the reduction and by the consequent increase in surface area available for reaction with hydrogen chloride. Can the author indicate whether any work has been done along these lines, and whether it has been established that the reaction depends on the surface area available? The second point relates to the application of this kinetic study to the actual operation of a Torco reactor. It was once planned to feed the sodium chloride, together with the reducing agent, into the top of the segregation chamber. In the paper, the author mentions that it has now been established that the segregation chamber behaves, to all intents and purposes, as a fluidized bed, and that there is, therefore, a rapid evolution of gas in the lower regions of the chamber, which, it is imagined, displaces the gas phase rapidly. Since the reaction between sodium chloride, water vapour, and the aluminium silicates in the ore is extremely rapid, one wonders how much of the hydrogen chloride produced is removed from the reaction zone before it has had time to react with the copper minerals. The extremely low consumption of sodium chloride (which is a vital feature of the Torco process) must depend on an extremely rapid circulation of the hydrogen chloride gas to all the ore particles in the segregation chamber. One wonders, therefore, if a system for the introduction of the sodium chloride into the bottom regions of the chamber might not result in even greater efficiency in the utilization of sodium chloride. ProC D. D. Howat (Fellow): All of us who have been concerned with the study of chemical reactions at high temperatures are keenly interested in kinetics and are well aware that this is not an easy study experimentally. Dr. Brittan is to be congratulated on the development of neat experimental methods and for his full discussion of the results obtained. Although the segregation process for the extraction of copper from oxide and silicate ores has been known for almost fifty years, the fundamental chemical and physical changes involved have been little understood and the fundamental data are very scanty. The work now in progress at A.A.R.L., together with that sponsored by the Anglo American Corporation in other research institutions throughout the world, is bound to produce new fundamental data and a much more complete understanding of this rather fascinating process. It is already apparent that some of the old and well-worn chemical reactions that were postulated to occur, just cannot take place in the way which was formerly accepted. Dr Brittan's work, carefully conducted and thoroughly analyzed as it has been, still leaves us with one great outstanding problem. The thermodynamic data and the possible reactions set out in Table I (page 281) of his paper leave us asking, in complete despair, how can copper be converted into a volatile chloride in the presence of HCl, CO and carbon at temperatures about 800°C? The thermodynamics all combine to show that copper should be reduced to metal as the first step in the process. This brings us right up against the second problem. If copper were reduced to the metal how would HCI convert it to the volatile chloride? On top of these problems is the unknown reason for the very high speed of reaction between CO, HCI and the ground copper ore. Still further into the region of the unknown is the reaction by which gaseous hydrochloric acid is produced in the actual process. Perhaps Dr Brittan is feeling grateful that he doesn't have to try to explain this reaction-at this stage of the research programme at least. The results very clearly show that both CO and HCl gas are essential for rapid production of the volatile copper chloride. Dr Brittan states that 18 minutes were required to attain 83 per cent extraction with HCl gas alone and this was reduced to 4 minutes when CO was

Jan 2, 1970

By M. Reuter, U. Boin

The target of the current metallurgical industry is to recycle and utilize all their by-products, so as to close the sustainable production loop. Slags are the important wastes and by-products of metallurgical industry, which have been treated, recycled and used worldwide. The present paper summarizes the current status of utilization of the various slags from ferrous and non-ferrous metal production, as well as waste incineration, and recycling of salt fluxes in secondary metal production. In addition, the environmental issues of the metallurgical slags are addressed. The metallurgical oxide slags have stone-like properties and, thus, their major applications are in civil engineering field. The slags should be recycled, modified and processed in a proper way, by taking the environmental impact into consideration. With the treatment of salt slags for melting aluminium scraps, as an example, the recycling and elution properties of the salt slags are discussed. Our research indicates that the viscosity of the salt flux is increased with addition of non-metallic components, that a certain percentage of fines remain in suspension, which determines the viscosity, affecting the settling of heavier materials. High slag viscosity will lead to more fine aluminium metal entrapped in the salt slag (may also be seen as a high temperature slurry), and thus increase the load of salt slag recycling. Keywords: recycling, environment, metallurgy, slags, fluxes, salts.

Jan 1, 2004

By I. Ewart, E. Roux

Skorpion Zinc operates a solvent extraction/electrowinning zinc plant producing over 150 000 t/a of special high grade (SHG) zinc (>99.995% Zn). In the current flow sheet copper and cadmium are cemented from the neutral leach solution using zinc powder, to prevent galvanic corrosion of plant equipment and build-up of copper in plant solutions. This process is effective but unfortunately consumes expensive zinc dust and produces a low value copper product. The EMEW® (Electrometals electrowinning) technology of Electrometals Technologies Ltd. uses a rapid flow of electrolyte through a cylindrical cell to deposit copper or other metals from dilute solution (less than 1 g/L) [1]. The copper can be deposited as powder or may be harvested as a tubular cathode product. The advantage of the use of EMEW® technology at Skorpion Zinc will be to eliminate the use of expensive zinc dust and to produce a high value copper product for sale. Following extensive bench test work, a piloting campaign was carried out at Skorpion Zinc. The tests demonstrated the ability of the EMEW® cell to effectively remove copper from solution through direct electrowinning at lower operating cost than cementation. Keywords: Skorpion Zinc, copper, cementation, direct electrowinning, EMEW®

Jan 1, 2007

By T. A. Muhlare

"Large quantities of slag are produced every year and dumped on slag heaps around the world, posing a potential environmental threat due to contained heavy metals content. Slags are waste minerals produced during furnace smelting of minerals. The contained values of base metals may be released into the environment under natural weathering conditions polluting ground and underground water. Copper losses into slag are a continuous problem in pyrometallurgical smelting operations. Due to shortage of ore concentrate, various mining operations have initiated the recovery of copper from dumped slag. The typical process uses crushing, milling and flotation. These operations create excessive liner wear rates during crushing and milling because the bond work index of copper slag is 22 kWh/t and this is considered a very hard slag. The process also has reduced flotation recoveries and requires high flotation reagent consumption. The present work shows that digestion of the milled copper slag slurry with concentrated sulphuric acid at a temperature of 250oC forms a solid cake and leaching of the solid cake with water is a possible option. However, the consumption of sulphuric acid can have a considerable cost on the recovery process. This can be limited by producing a leach residue containing gypsum and silicate materials as amorphous silica which are suitable for use as an additive for Portland cement. Laboratory results show that 90% copper recoveries can be achieved. The leaching is conducted to prevent silica gel formation and iron co-extraction. Leaching is optimum if silica gel-free solutions and good filterable residue are obtained. INTRODUCTION Slag is a waste by-product of smelting and converting operations in metallurgical plants. Both smelting and converting operations are aimed at iron separation from base metal sulphides in non-ferrous metallurgy. The iron sulphide component of the ore or concentrate is oxidized to iron oxide at higher temperatures in the molten form, and then fluxed with silica and other oxide rock minerals to produce fayalite (2FeO•SiO2) slag (Schlesinger et al., 2011). Slags are generally deposited on dumps and considered to be unreactive materials. Smelter copper slag contains significant concentrations of several potentially toxic elements, including arsenic, lead, cadmium, barium, zinc and copper. These elements can be released into the environment under natural weathering conditions and cause pollution of soils, surface waters and groundwater (David, 1998)."

Jan 1, 2015

By P. J. Lloyd

Society is fully justified in its concern about the environment, but that concern may, at present, be too uncritical It is, for instance, essential to establish that there is a valid environmental impact, and that there is truly a causal link between a human activity and an observed shift in environmental conditions. In establishing these, it inessential to recognize that scientific proof offers consistency, but that it can never provide completeness. Thus, no amount of scientific endeavour can ever prove that mitigation of the effects of human activity will be perfect. Society normally expresses its concern through legislation. The existing legislation in South Africa is reviewed, and it is found that [>legislation regarding water is probably too stringent to be effective, >legislation regarding air and land use is probably too lax, >all the environmental legislation is too fragmented to be efficient, and the Department of Environmental Affairs is hamstrung by inter-departmental rivalry. ] It is accordingly concluded that, if means are to be found for ensuring adequate mitigation of environmental impact in the face of ongoing economic development, it is necessary to establish a Judicial Commission on Environmental Affairs advised by, but given greater powers than, the departments currently active in environmental maters. This Commission should be charged with creating, perhaps over a decade, a body of legislation that would guide development while protecting the environment in its widest sense.

Jan 1, 1994

By R. Altindag

The rock brittleness is one of the most important rock properties that affect the drill-ability of rocks. It is supposed that the increase in rock brittleness causes the increase in penetration rate. The drill-ability of rocks depends on many factors. These factors can be separated as two parts depending on a machine, bit type, rotational speed, thrust, blow frequency, flushing, and rock properties, brittleness, compressive strength, structure, hardness, etc. In this paper, the relationship between brittleness and drill-ability index, which is used in the prediction analysis of the penetration rate of rotary blast hole drills, were examined using regression analysis. It is also seen that significant correlations exist between brittleness and the drill-ability index.

Jan 1, 2002

By J. D. Steenkamp

In a paper presented at the Heavy Minerals Conference 2001 a laboratory scale investigation on the beneficiation of an ilmenite containing waste stream by magnetizing roasting was reported1.This waste stream contained a high concentration of ilmenite, which was contaminated by chromite and other gangue minerals. Interpretation of the results of the investigation was based on the assumption that during magnetizing roasting of an ilmenite concentrate, under the oxidizing conditions, the magnetic properties of the chromite would remain constant. In this article thefollowing hypothesis was tested: The magnetic susceptibility of chromite would remain constant during magnetizing roasting of an ilmenite concentrate under the oxidizing conditions reported by Gouws and Van Dyk1. The results of this study indicated that the hypothesis is not true. The increase in the magnetic susceptibility is not significant though and it is expected that it will not influence the final ilmenite quality to a large extent. Based on the assumptions that the bulk magnetic susceptibility of the waste stream samples represents that of the ilmenite in the waste stream, and the bulk magnetic susceptibility of UG1 chromite that of the of chromite in the waste stream, the results also served to confirm that the waste stream with a high chromite content should be roasted under oxidizing conditions, in a reactor with a well-defined retention time distribution (i.e. a fluidized bed reactor), at a roasting temperature of 750°C (rather than the higher temperature ranges of 800°C and 850°C).

Jan 1, 2003

By Willem Le Roux

Occupational health and safety is regulated by statute and common law. South Africa has, like many other countries, legislated extensively on occupational health and safety. Despite the body of legislation, the common law continues to play an important role as far as it concerns contractual, delictual and criminal liability. Two major Acts regulate occupational health and safety viz. the Occupational Health and Safety Act, No. 85 of 1993 (?OHASA?) and the Mine Health and Safety Act, No. 29 of 1996 (?MHSA?). The MHSA applies to mines and works as defined and related aspects. OHASA applies to other industries but does not apply to employers and workplaces to which the MHSA and certain matters covered by the Merchant Shipping Act, No. 57 of 1951 apply. See section 1(3) of OHASA. 2 The employer?s health and safety obligations in terms of the MHSA 2.1 Section 2(1) of the MHSA provides amongst others, as follows: ?The employer of every mine that is being worked must: (a) ensure, as far as reasonably practicable, that the mine is designed, constructed and equipped: (i) to provide conditions for safe operation and a healthy working environment (ii) ... (b) ensure, as far as reasonably practicable, that the mine is commissioned, operated, maintained and decommissioned in such away that employees can perform their work without endangering the health and safety of themselves or of any other person?.

Jan 1, 2011

By T. van Heerden

A range of supported gold catalysts was prepared by ion exchange, varying many of the preparation variables, including concentration in the precursor solution, washing procedure, as well as drying and calcination procedures. These catalysts have been characterized extensively. TEM images show essentially the same crystallite size distributions, between 2?5 nm, for almost all catalysts, the only exception being catalysts not washed in ammonia, which did not show any small crystallites. Additional characterization with SEM yielded an interesting discovery. Catalysts that appear identical on the TEM also contain some large crystallites in the range of 50?500 nm. Differences in dispersion due to the drying procedure not seen on the TEM can now be observed. Oxygen chemisorption is being investigated as an additional method to characterize gold based catalysts to complement the typically used electron microscope techniques.

Jan 1, 2012