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1. Journals Marine Geotechnology, edited by R.C. Chaney. London, Taylor & Francis (Rankine Road, Basingstoke, Hants RG 24 OPR, OK), 1988. vol. 8, $50. Published quarterly. Mineral Resources Engineering, edited by C.T. Shaw. London, Taylor & Francis (address as above), 1988. vol. 1, $99. Published quarterly. 2. New publication World mining map of non-ferous metals (with 4 regional maps and a booklet). Reinhard Ryborsch (Postfach 2105, D-6053 Obertshausen bei Frankfurt am Main, West Germany). International edition: bilingual, English and German. 7 colour-print: Size (unfolded) 138 X 100 cm. US$ 19 (excl. mailing) for folded map with hard cover (Order No. 61-4); US$ 25 (excl. mailing) for unfolded (flat) map with hanging strip (Order No. 62-2). 3. Mintek reports Report M326 Crevice corosion and other localized corosion behaviour of 3CR12 corosion-resisting steel in synthetic minewaters, by D. Howarth. May 1988. 28 pp.Report M332 The carbon-regeneration furnaces at Mintek, by P.M. Cole and P.J. van Staden. Nov. 1987. 31 pp. Report M344 The flotation of pyrite with amine colectors, by R.D. Hil, E.W. Giesekke, and P.J. Haris. Mar.'1988. 25 pp.Report M345 The substitution for chromium in steels: Progress and trends, by M.B. Cortie. Mar. 1988. 14 pp.Report M346 The direct reduction of sulphur dioxide to elementalsulphur, by I. Henderson. Mar. 1988. 30 pp.Report M349 The optimization of a process using weak-base resin for the recovery of gold from plant solutions and pulps, by A. Mehmet. Apr. 1988. 17 pp. Report M353 The determination, by ion-interaction chromatography, of sulphur species in cyanide solutions, by Pohlandt-Watson, M.J. Hemmings, D.E. Barnes, and a.w. PansL Jun. 1988. 7 pp. Report M355 A laboratory-scale continuous-feed resistance furnace, by K.P.D. Pery, A.S.E. Kleyenstiber, andC.T. Logan. May 1988. 9 pp. Report M357 The determination of minor and trace elements in activated charcoal, by RT. Eddy and G.J. Wal. Jun. 1988. 14 pp. Report M358 The sensitive determination of germanium by atomicabsorption spectrophotometry using electrothermal atomization, by G.D. Marshal. Jun. 1988. 11 pp. Report M92D The recovery of tin and tungsten from various zones of the Van Roois Vley deposit, including setling and filtration, by C.F.B. Coetzee. First issued Apr. 1983; reissued Apr. 1988. 40 pp. World Environment Day Symposium* Mintek auditorium in Randburg was the venue on Thursday, 2nd June, 1988, for more than 200 interested delegates at the annual World Environment Day Symposium, organized jointly by the Federation of Societies of Professional Engineers (FSPE), The Environmental Planning Professions Inter-disciplinary Committee (EPPIC), The Habitat Council, and The Associated Scientific and Technical Societies of South Africa (AS&TS). The theme 'Cultural and Historical Conservation' was adapted from the theme of the Department of Environment Affairs for the 1988 Environment Week.

Jan 1, 1988

By S. L. Crouch

This paper describes a numerical procedure for the computation of the displacements and stresses induced by the mining of a tabular ore deposit in faulted ground. The numerical procedure is based on the displacement discontinuity method, which can be regarded as an application of Salamon's 'face element principle' to cases in which the reef plane is faulted into several sections, or even folded in an arbitrary way. An element of a fault is modelled in the same manner as an element of the reef, except that a fault element is allowed to undergo permanent inelastic deformation (i.e., slip) if the stresses transmitted through the fault-filling material satisfy a Mohr-Coulomb yield condition. An iterative technique for the incorporation of a Mohr-Coulomb yield condition in the general displacement discontinuity method is explained. This technique is illustrated by means of a simple example involving the mining of a flat-lying reef intersected by a fault. The displacements and stresses at selected locations along the fault are examined as the reef is mined, element by element. A consideration of the energy changes reveals that the occurrence of slip along a fault leads to a sudden increase in the energy release rate due to mining as the excavation intersects the fault. The energy release rate then falls off abruptly to a value less than that which would occur in the absence of any slip, and gradually builds up to re-establish the conditions that would exist in unfaulted ground.

Jan 1, 1979

By Isobel Clark

This paper reviews some of the traditional methods of reserves estimation based on lognormal distribution with or without an additive constant. In the rush to computerize mine planning and grade control generaly, proven estimation methods, such as Sichel's t, seem to have been overlooked. By use of a personal microcomputer, the traditional methods were investigated thoroughly, the results being tested against existing tables and figures. The fitting of a lognormal distribution, including estimation of the additive constant, is discussed, and the calculation of Sichel's t estimator and its associated confidence limits is described in detail. Payability calculations for a lognormal distribution are also covered briefly. The advantages of personal computers, particularly speed and ease of use, are emphasized. This low-cost approach permits detailed investigation of some of the assumptions and approximations inherent in the established methods of calculation. Several disturbing discrepancies are revealed in various factors, and some generalized statements made by previous authors are found to be over-simplistic. Full details are given of acceptable approximations and computer algorithms, and concern is expressed over the possible loss of proven methods under the welter of new, sophisticated computer software.

Jan 1, 1987

By T. Oberthür

The Great Dyke of Zimbabwe constitutes the world&apos;s second largest reserve of platinum group elements (PGE) after the Bushveld Complex in neighbouring South Africa. Within the Great Dyke, economic concentrations of PGE are restricted to sulphide disseminations of the Main Sulphide Zone (MSZ), which are currently mined at the Ngezi, Unki, and Mimosa mines. Near-surface oxidized MSZ ores have a large potential. Their total resources are in the range of 160-250 Mt; however, all previous attempts to extract the PGE from this ore type have proved uneconomic due to low PGE recoveries (<< 50 per cent) achieved by conventional metallurgical methods. Within the ores of pristine, sulphide-bearing MSZ, the PGE are bimodally distributed. Platinum occurs mainly in the form of discrete platinum group mineral (PGM) grains (mainly bismuthotellurides, sulphides, and arsenides), whereas approximately 80 per cent of the Pd (and some Rh) is hosted in pentlandite. Within the oxidized MSZ ores, the PGE are polymodally distributed. Whereas the arsenide- and sulphide-PGMs that make up approximately 25 per cent of the original Pt content of the ore largely remain stable (relict PGMs), the remaining PGMs are disintegrated. The base metal sulphides are destroyed, partly releasing their base metal and PGE contents, and are replaced by iron oxides or hydroxides. Unspecified amounts of the PGE are redistributed and either form secondary PGMs, are found in chemically and mineralogically ill-defined (Pt/Pd) -oxides or hydroxides, or in iron-hydroxides, Mn-Co-hydroxides, and in secondary silicates. The problematic processing of oxidized MSZ ores is attributable to their complex nature and polymodal distribution of the PGE, prohibiting a significant upgrading of the ores by conventional metallurgical methods. Therefore, only bulk leaching methods are viable for ore treatment, and novel metallurgical methods have to be developed for the processing of these ores. Our ongoing work aims at locating the PGE in their mineralogical form in order to understand the mineralogical balance of the PGE in the ores and thereby facilitate the evaluation of metallurgical options for their recovery. A short overview on options and recent advances regarding the recovery of the PGE from oxidized ores is given.

Jan 1, 2013

By A. M. Croll

The efforts to mechanize narrow reef tabular orebodies unique to this country, are not new. They began in the early 1900s and have over the last century seen varied rates of progress and success. Efforts to mechanize with the use of scraper winches can be regarded as a success. The scraper winch replaced the hand lashing of ore and development waste into hoppers. The initiatives by Anglo Gold and the then JCI to implement trackless mining equipment into these environments had limited success but in general were looked upon as a failure?at least this is the perception. The Chamber of Mines Research Organization (COMRO) in conjunction with various mining houses tried to make a success of the impact ripper. Diamond wire cutting of the stope face is another, which has been tried and tested. The use of high-pressure water to cut rock has also been trialed. These attempts, amongst others, are just a few of the efforts that have not been successful in the quest to make life at the face safer, more productive and easier for our employees. Against the backdrop of an emerging South Africa, the associated high expectations of historically disadvantaged people within the country, a strong rand, demands for better safety, lower cost of production and more efficient operations, mechanization has the chance to make a difference to Anglo Platinum?s business. Mechanization has certainly changed the way other commodities such as coal, base metals and diamonds are won from the earth. Mining?s image is still one of people crawling around in stopes?a dirty, ?low-tech? business. Anglo Platinum?s latest journey to mechanize its underground operations began in earnest in1998 and has enjoyed some successes and failures. The opportunity to change the way mining is done within the company is there. Technologies are evolving, and the implementation of mechanization can certainly assist with the achievement of the goals referred to above. When mechanization commenced, targets were set for the company to achieve but these have been revised in the light of recent events relating to the strengthening rand and a slow-down in expansion plans. However, the intent to mechanize appropriately in Anglo Platinum still exists.

Jan 1, 2004

1. Book review The geology of some northern Nigerian anorogenic ring complexes, by J.N. Bennet, D.C. Turner, E.C. Ike, and P. Bowden. The British Geological Survey: Overseas Geology and Mineral Resources, no. 61. London, HMSO, 1984. 66 pp. Reviewer: R.H. Ingram 2. Mintek reports Report M226 The separation of tungsten and molybdenum by the formation of sulphide complexes and extraction into a weak-base resin. Report M228 The solvent extraction of zinc, iron, and indium from chloride solutions by neutral organophosphorus compounds. Report M230 Shear flocculation of fines for improved flotation. Report M239 A mineralogical study of residue from the column leaching of a low-grade nickel-copper sulphide ore.3. Recent publication Driger Review, special edition, act. 1985. 48 pp. (Copies obtainable from Roly Nyman, HAMPO PMIE Division, P.O. Box 170, Joham1esburg 2000.) Black Mountain&apos;s achievement The 13 689 t consignment of metal concentrate that was loaded aboard the bulk carrier &apos;George&apos; and shipped out of Saldanha Bay early in December brought the number of tons exported from Black Mountain via the Sishen-Saldanha railway and harbour system to one million.

Jan 1, 1986

By E. Pothas, P. R. Jochens, L. B. McRae, D. D. Howat

Discussion H. J. S. Kriek*: During 1956 Iscor became interested in the acid smelting in a blast furnace of siliceous ores containing 15 per cent silica. This interest was as a result of the successful acid smelting in Germany at a CaO/SiO2-ratio of between 0.4 and 0.8. Acid smelting can become an economic proposition due to a reduction in slag volume accompanied by a lowering of the coke rate. Paschke and Hannel1 found that a free flowing slag (i.e. viscosities of less than two poise) would not be obtained at C/S-ratios lower than 0.82 and that the addition of 3 to 4 per cent alkalies became necessary. In titanium bearing slags it was established that maximum fluidity is obtained with 10 per cent TiO2. Acid smelting resulted in increased sulphur pickup with decreasing O/S-ratio and desulphurization with soda-ash outside the blast furnace was necessary. At Iscor Mr B. B. Segal established that, for slags with (CaO+MgO)/SiO2-ratios of 0.5, 0.6 and 0.7, the addition of alkalies and TiO2 resulted in a substantial decrease in melting point and also of viscosity. At a (CaO+ MgO)/ ratio of 0.53 the addition of 10 per cent TiO2 is as effective as the addition of 3.75 per cent alkalies. This means that a blast furnace can be run without the addition of alkalies and consequently the detrimental effect of alkalies on the lining can be avoided; also that fluid slags with 5 to 15 per cent MgO and 8 to 10 per cent Al203 can be obtained with 10 per cent TiO present. (CaO+ MgO)/SiO2-ratios smaller than 0.5 the addition of titania resulted in a decrease in melting point and also in an increase in viscosity. In ceramic crucibles the slag did not become more viscous with time but thickened appreciably when held in graphite crucibles. No titanium carbonitride could be detected and only Ti20a was found. Prof Howat also did not obtain the TiC-TiN solid solution and therefore it appears that the thickening of the slag is not solely due to the formation of the titanium carbonitrides. This is in agreement with the work of Michailov and Belyakova2 who found that TiO2 decreased the viscosity of acid slags and increased the viscosity of basic slags; that titanium sesquioxide (Ti203) and titanium monoxide (TiO) increased the viscosity of both acid and basic slags and that titanium carbide, which is found in larger amounts in basic slags than in acid slags, considerably increases the viscosity of both acid and basic slags. Patzak et al3 describes the following oxygen deficient titanias, viz Ti2O, TiO, Ti203, Ti3O5, Ti4O7, Ti5O9, Ti6O11- Ti10O9,Ti7O13,Ti8O15 and Ti9O17. Until about 1959 a fair amount of data was available on the viscosity of normal slags to which titania was added2, 4, 5, 6, 7, 8, 9. At Iscor an electromagnetic viscometer based on the viscometer developed by Bockris10 was built. Progress was very slow as our main attention was centred around problems associated with the operation of the Rotor process. During 1962 and 1963 two papers containing fundamental work on the viscosities of slags in the systems CaO-SiO2-TiO211 and CaO-SiO2-TiO2 at 0 per cent A1203, 10 per cent Al203 and 20 per cent Al20312 appeared. The work of Ross and Ohno obviated the necessity of any further work by Iscor on the viscosity of blast furnace slags containing titania. Ross gave the following optimum compositions for blast furnace slags containing titania. Titania % Lime % Silica % Lime/Silica-ratio 20 30-42 38-50 0.6-1.10 30 27-36 34-43 0.63-1.06 40 24-32 28-36 0.66-1.14 The optimum temperature is 1,500°C because at this temperature the largest workable composition range exists. Higher temperatures result in narrower optimum composition ranges and in increased rates of thickening. The factors affecting the rate of thickening were also investigated by Ross and Ohno, as the main problem with slags containing titania is not finding compositions which are fluid at operating temperatures but the thickening due to a lack of oxygen. With these slags when there is insufficient oxygen to co-ordinate each silicon atom with its own individual shell of four oxygen ions, some sharing of oxygen ions must occur, resulting in the formation of large silicate ions and a rise in viscosity. The formation of oxygen deficient titania can be prevented as long as iron oxide or manganese oxide remain in the liquid slag. Iwase13 and also Viens, Camp bell and Rogers14 obtained in the experimental smelting of ilmenite ores fluid slags with more than 3 to 5 per cent manganese oxide and ferrous oxide in the slags. It appears that titaniferous ores can be smelted in blast furnaces under the following conditions: (1) Acid slag operation, i.e. at C/S-ratios from O.6-1.0. (2) Dolomite can be used as a flux as MgO is a network modifier. (3) The Al2O3-content must be less than 15 per cent as alumina promotes thickening of the slag and decreases the optimum composition range. (4) The hearth temperature must not exceed 1,500°C. According to Ross a definite amount of energy is required for each ton of iron produced and in order to increase production an increase in heat

Jan 6, 1969

1. Reviews Extraction metallurgy, by J.D. Gilchrist. 3rd ed. Oxford, Pergamon, 1989. 431 pp. £50.00 (hard cover), £14.95 (soft cover). Reviewer. Prof. M.D. Brayshaw. Reviewer. Prof. M.D. Brayshaw This third edition bears testimony to a work that contains much of value for lecturers, students, and practising metallurgists. Mineral processing in the United Kingdom, edited by P.A. Dowd. London, Institution of Mining and Metallurgy, 1988. 223 pp. £25.00. Reviewer: Prof. M.D. Brayshaw It came as both a pleasure and a challenge for a resident of South Africa (with its volume and diversity of minerals) to review a book focusing on the beneficiation of ores from the sparsely endowed mineral chests of the United Kingdom. 2. Recent issues Rock mechanics. Theory and applications with case histories, by W. Wittke. Translated from German by R. Sykes. Berlin, Springer-Verlag, 1990. 1076pp. DM248. McGregor&apos;s quick reference to the JSE. Pretoria, Juta & Co., 1990. R39,95. McGregor&apos;s economic alternatives. Pretoria, Juta & Co., 1990. 400 pp. R34,95. 3. New journals Journal of Materials Chemistry, edited by A.R. West and J. Leader, and published two-monthly by The Royal Society of Chemistry (Blackhorse Road, Letchworth, Herts SG6 IHN, OK). £195 per year, starting in 1991. New Materials World, published monthly by Lawrence Cooklin (15 Selvage Lane, Mill Hill, London NW? 3SS, OK). £180 per year. Mendeleev Communications 4. Books from The Institute of Metals The following were published recently, or are to be published soon, by The Institute of Metals, 1 Carlton House Terrace, London SWl Y SOB, OK. Illustrated case histories of marine corrosion. 1990. Modern steelmaking methods, by C. Moore and R.I. Marshall. 2nd ed. 1990. Grain boundary fractures, by S.F. Pugh. 1991. 160 pp.

Jan 1, 1991

By F. M. Russell

Scheduling the extraction and rates of production for an underground mining operation is a common task carried out during the life of a mine, from the initial feasibility study through to the final production phase. The scheduling process can be complex when several requirements have to be satisfied, for example, match mill capacity and maximize net present value while constrained by mining sequence and availability of resources. Many mines have computer programs relevant to their operation but because they are site-specific they need considerable reprogramming to be adapted to other operations. Other methods currently in use also suffer drawbacks: manual spreadsheets are laborious, computerized spreadsheets cannot handle sufficient constraints, and operations research techniques require a complex mathematical model to be set up. The use of network analysis techniques has been proposed previously but the unavailability, until recently, of easily used computer programs has not allowed this approach to be developed. There are now many programs which can be used; however, the particular program which has been adapted to mine scheduling is Metier&apos;s &apos;Artemis-PC&apos; running on IBM-XT or compatible machine. The use of network analysis offers many advantages, including the use of a logical mining sequence; activities can be as detailed as required; related activities can be readily incorporated, and costs can be easily extracted. A heuristic approach is necessary to optimize the schedule, but the ease and speed of program execution enables this to be readily achieved, while leaving the mining engineer in full control. The paper discusses the problems encountered in mine scheduling and current solutions. It then details the use of network analysis techniques and the application of the Artemis-PC program to solve mine scheduling problems.

Jan 1, 1987

1. Book review Economics of mineral exploration in Australia, by Brian W. Mackenzie and Michel L. Bilodeau. Australian Mineral Foundation (Private Bag 97, Glenside, South Australia 5065). 198pp. $A58. $A45 for students. Reviewer: D.G. Krige 2. Recent publications Combustion of coal liquid mixtures, by G.F. Morrison. London, lEA Coal Research, 1983, 52 pp. Hafnium: physico-chemical properties of its compounds and alloys. Vienna, International Atomic Energy Agency, 1981. 408 pp. 620 Austrian schillings. Canadian mines handbook 1984-85, edited by CD. Gardiner. Toronto, The Northern Miner, 1984. 500 pp. $24. Drilling equipment in Canadian mining and exploration- opportunities for the industry, by AJ. Jakubek. Toronto, Micromedia Limited (144 Front Street, Toronto M5J IG2, Canada), 1984. Research report 1984. Port Melbourne, Division Mineral Chemistry, Commonwealth Scientific and Industrial Research Organization, 1984. 190 pp. Surface engineering Wear is an unavoidable problem in virtually all products of the engineering industry, from the minute bearings of instruments to the cutting surfaces of earth-moving plant. Safety in mines About 50 papers will be presented at the 21st International Conference of Safety in Mines Research Institutes at Sydney from 21st to 25th October, 1985. Some 250 delegates from 30 nations are expected to attend. Professional development seminars March 18th-21st, 1985 Underground Blast-hole Mining May Ist-3rd, 1985 Mineral Processing of Gold Ores May 6th-10th, 1985 Mineral Comminution Systems May 13th-15th, 1985 Processing of Metallurgical Residues Analytical chemistry The Second Symposium on Analytical Chemistry in the Mining, Exploration and Processing of Materials is to be held in Pretoria from 15th to 19th April, 1985.

Jan 1, 1985

By H. J. R. Way

Dr H. J. R. Way: In his reply to my queries Mr Bath still adheres to his statement that it may be possible to replace the estimated R200 million reduction in gold revenue in 1980 by exports of ferro-chromium, and apparently based his conclusion on a growth rate in demand for corrosion resistant steels of 10 per cent per annum in the next few years. As I pointed out at the previous discussion at Middelburg this would entail a production of about one million tons of ferro-chromium in 1980. Exact statistics of South African ferro-chromium production are not available, a fact which is very much to be regretted. But statistics of exports of ferro-chromium are available from "Foreign Trade Statistics" of the Customs and Excise. These show that South Africa exported 99,580 short tons in 1967. The quantity of ferro-chromium consumed in South Africa in 1967 was fairly small and we can therefore assume that the production was 99,580 tons. For this amount to grow to 1,000,000 tons in 1980 would require a growth rate of 19.4 per cent per annum. At 10 per cent this would grow to 344,000 tons and at 7 per cent 240,000 tons. The following growth rates in the form of average trends are pertinent: STEEL - Per annum World steel ingot production - 1957-67 – 6.8 per cent Free-world ingot production - 1957-67 – 6.0 per cent Consumptions of crude steel by four major Free-world Consumers - 1957-67 - 5.6 per cent CHROMITE World production - 1952-67 - 2.5 per cent South African total sales - 1948-67 – 5.4 per cent South African exports - 1948-67 – 3.9 per cent United States consumption, very irregular, but - 1962-67 – 6.5 per cent FERRO-CHROMIUM - Per annum United States consumption - 1952-67 - 6.5 per cent South African exports - 1957-62 - 25.2 per cent South African exports - 1962-66 - 41.9 per cent South African exports - 1966-68 - 21.2 per cent These figures show that exports of ferro-chromium have been growing at a very high rate varying from 21.2 per cent to 41.9 per cent since 1957, but in view of the market situation it seems somewhat doubtful if that high rate can be maintained up to 1980. Of our 1966 exports U.S.A. took 63 per cent which was 53 per cent of their total imports. Their average consumption is trending at a growth rate of 6.5 per cent per annum from 1952-67, but the growth from 1964-66 was 5- 2 per cent and 1966-67, -12 per cent!

Mar 1, 1968

By L. Hermansson

Trackless mining at LKAB, Sweden by L HERMANSSON (Visitor) Because the usual means of transporting men and material to the working sites was time-wasting and uneconomical, inclined subways were installed in the Kiirunavaara Mine that permitted the access of rubber-wheeled vehicles to all underground working sites. Men and material could now be transported direct to the working site with no reloading. An increase of 8 to 12 per cent in underground productivity has been obtained by the use of this new transportation system. Rubber-wheeled vehicles are also used in development, production drilling, and loading. In December 1958, the management of the Luossavaara Mine decided to cut the costs of waste-rock haulage. The system of waste transport was the conventional one of shuttle car, rail haulage, and skip hoisting. After consideration of other methods, it was decided to re-equip the mine completely for a system of trackless transport. Results show that mining costs per ton of ore were reduced by 27 per cent, and mining costs per ton of waste were reduced by a spectacular 80 per cent. The new system has proved to be more flexible, and ore grading has been simplified since there is no longer any mixing in ore passes and chutes.

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

Jan 1, 1978

By D. Duncanson, R. J. Davidson, C. G. S. Schmidt, G. A. Brown, J. D. Taylor, N. W. Hanf

The development of a cyanidation route for the treatment of gold-plant gravity concentrates is described. This work was carried out as a natural consequence of earlier improvements to the recovery of gold in the gravity sections of existing gold plants. Concern about the environmental aspects of the conventional amalgamation process was further motivation in the development of an alternative procedure. It was found that the partial pressure of oxygen in solution was the most critical reactionrate determining factor. The rate of gold dissolution was shown to be controlled by the reaction rate below approximately 25°C and by diffusion above that temperature. Thus, the efficiency of pulp agitation and oxygen partial pressure permitted a high degree of control over the rate of gold dissolution. The use of oxygen rather than air significantly increased the dissolution rate of both gold and silver. The solubilities of sodium, calcium, and potassium aurocyanide complexes under simulated reaction conditions were studied, and it was established that the values for the sodium and calcium complexes were limited, requiring the use of lower pulp densities and higher reaction temperatures for the treatment of very high-grade concentrates. The presence of excessive amounts of tramp iron in such concentrates of up to 25 per cent by mass was found to result in poor gold dissolution owing to the cementation of gold under certain reaction conditions. The maintenance of high oxygen partial pressures, together with high cyanide concentrations, high pH levels, and the presence of calcium in the system, all aided in passivating the tramp iron. Copper, gold, and iron were shown to be the main cyanide-consuming elements. Optimum gold dissolution was obtained in a closed bench-scale reactor incorporating a flotation-machine type of agitator for effective pulp agitation and oxygenation, the oxygen being recirculated. The reactor was designed to operate under a very small positive pressure. After reaction times of 2 to 6 hours at temperatures in the region of 30°C, both the gold and the silver dissolution averaged 99,S per cent. In the case of very high-grade concentrates, it was shown that gold could be floated in the same reactor by drawing off the flotation concentrate through a port in the reactor. A flotation recovery of 87,4 per cent of the gold in 3,9 per cent by mass was obtained. The recovery of osmiridium from cyanide-leach residues and the electrowinning of gold from the pregnant liquors were also considered as processing aspects in the treatment of gold-plant belt concentrates. Final table tailings with a gold concentration of 10to 12 gft were obtained. The results of the investigation strongly suggested that an alternative cyanidation route for the treatment of gold-plant concentrates was technically feasible, provided the system included effective agitation and oxygenation. In the case of lower-grade concentrates, cyanidation followed by the electrowinning of gold appears to be the simplest route, while, for high-grade concentrates, flotation and smelting of the concentrate followed by cyanidation of the flotation tailings are suggested. The use of oxygen rather than air for sparging the pulp would depend largely on the grade of material to be treated. A preliminary comparison of the costs related to the amalgamation and cyanidation processes for the recovery of gold from such concentrates indicated very similar capital and operating costs. It was concluded that other considerations, e.g. toxicity, safety, and lower labour requirements, were more important than such small cost differences. Furthermore, the &apos;opening&apos; of the gold-plant gravity circuit, which would be made possible by a cyanidation route, would further reduce the operating costs because of the subsequent decrease in the tonnage to be treated.

Jan 1, 1978

By Z. Leng

"After detonation of the explosive, the final blast-induced damage area around the blast-hole falls into three categories, namely crushed zone, fractured zone, and elastic deformation zone. This paper presents a modified model to calculate the size of crushed zone around a blast-hole in drill-and-blast, with the hoop compressive stress and cavity expansion effect taken into account. The material in the crushed zone was assumed to be a granular medium without cohesion, but with internal friction. On this basis, the formula for crushing zone radius was derived. The proposed approach was verified with tests reported in the literature and the simulated results from SPH-FEM coupled models. A full statistical analysis was also carried out, and the predicted values were found to be in better agreement with the test results compared with other models. A sensitivity analysis of the modified model showed that the size of the blast-induced damage area is mainly affected by the following factors: rock mass properties, in situ stress, and borehole pressure. In particular, the roles of borehole pressure and in situ stress are discussed. Finally, suggestions are made on how the size of the crushing zones can be decreased. IntroductionOver the past decades, drill-and-blast has become the most commonly used technology in rock excavation. It is well known that in rock mass fragmentation with explosives, the annular rocks around the blast-hole are converted into fines. The formation of these fines consumes a significant part of the energy of the detonation, which in general is ignored in the determination of the efficiency of detonation (Glatolenkov and Ivanov, 1992; Furtney et al., 2012). Many studies show that only 20–30% of the total explosive energy is effectively used in fragmenting the rock, and up to 50% of the energy generated by conventional charges is wasted in overcrushing of the crushed zone and the inner part of the fractured zone (Ouchterlony et al., 2004; Sanchidrian et al., 2007). How to control the crushed zone to enhance the effective utilization of explosive energy, reducing the unit explosive consumption and the engineering cost, is therefore of great significance."

Jan 1, 2016

The proceedings of the Centenary Banquet that was held on Friday, 25 March 1994 at The Bal Room of The Carlton at 19:30 follow. Dr H. Scot-Russell The Chemical and Metallurgical Society of South Africa-a forerunner of this Institute-was founded at a meeting held in Johannesburg on the 24th of March, 1894. One of the traditions established early in the life of the Society was the Annual Dinner, which became a major event in Johannesburg&apos;s social calendar. It was always attended by people of stature in the industry and in the community, and was usually addressed by a high-ranking Government figure. We are delighted that, on this one hundredth birthday of the Institute, we have been able to continue the tradition. I welcome you, Mr State President, and your charming wife, and we look forward to hearing your address later this evening.

Jan 1, 1994

By D. G. Hulbert, I. K. Craig, S. P. Moult

This paper describes the application of a throughput-optimization strategy to a run-of-mine (ROM) mining circuit that is controlled by a multivariable controller. The objective of the optimization strategy is to maximize the ROM throughput at a specific product size. Results from long-term tests, which indicated how &apos;steady-state&apos; optimization can be achieved, were used for the throughput optimization. This strategy increased the ROM throughput by 11,4 per cent. A mathematical relationship relating residue to grind can be used to calculate the economic benefits of stabilizing control. It was estimated that the throughput optimizer could result in an increase in gold production worth Rl,4 million per annum when waste rock is treated, and that stabilizing control could increase the treatment plants annual revenue by R150 000. If the extra milling capacity brought about by the optimizer is interpreted as a capital cost saving of 10 per cent on the price of a new mil, then a capital cost saving of RI million could result.

Jan 1, 1992

By Neal Harries

The monitoring of rock slopes is a common procedure in the management of hazards associated with slope instability. Slope stability radar has become widely accepted as the state-of-the-art method for management of hazards associated with unstable slopes in open pit mines. This is confirmed by the rapid adoption of the technology since inception in mid 2003, with over 70 GroundProbe slope stability radars now in use at open pit operations throughout the World. The use of a monitoring system such as slope radar to identify slope hazards is only part of the solution to minimise the consequences associated with slope instability. The presence of the hazard must also be effectively communicated to the mine personnel, so that the correct response to the hazard can be enacted, and the risk mitigated to an acceptable level. In some surface mines, due to the geotechnical characteristics of the rock mass, hazards can develop rapidly; with slopes progressing from stability to ultimate collapse in just hours. In these situations, it is vital that efficient systems are developed to evacuate the area of the pit affected. Such a system has been developed by GroundProbe in South Africa, with the vital driving force for the project being provided by the Anglo Platinum Mogalakwena mine operation based at Mokopane. The mine based development of the system was perceived as an essential aspect in the development of the successful system. From an early stage, the mine provided essential input on both hardware and software developments to allow the system to be rapidly integrated into the mining operation. This paper describes the pit evacuation alarm system that was developed and discussed the benefits obtained from a mine based development program.

Jan 1, 2008

By Z. Saw, C. Poveda, N. L. Mai

Optimising the throughput of a mining network comprising all equipment from shovels to semi-autogenous grinding (SAG) mills is a complicated task, yet the potential benefits can be significant. Key modelling challenges stem from the sheer number of possible routes which material can travel, different equipment operating under different constraints, and the rate of SAG mills being calculated through a nonlinear function of the blended materials’ rock attributes. Nevertheless, this is an ideal onapplication area of mathematical programming, a branch of operations research. This paper describes an application of mixed integer linear programming (MILP) to optimise the mining equipment network of one of the world’s largest copper mines.The model takes as input a year-long monthly mine schedule of shovels in different pits, pushbacks, sequences and faces which produce material to be moved through a network of crushers, conveyors, silos, distributors, stockpiles and SAG mills, ultimately outputting a daily operational plan for all equipment that maximises the throughput of the network. All operational constraints and business rules are considered, including three quality attributes of copper ore. Comparing model results with actual data shows an improvement of 4.47% in network throughput tonnage. Additionally, although the model is solved using a free, open-source solver, optimal solutions are still achieved within a practical timeframe.

Sep 1, 2021

By C. S. Northcote, I. F. Buchan, L. F. Baars

In 1969 the Coal Division of the Anglo American Corporation was awarded the tender to supply coal to the new 3000 MW power station to be erected at Kriel, which is situated midway between the towns of Ogies and Bethal in the eastern Transvaal. The coal was to be supplied from the No. 4 Seam in the Kriel Area of the Witbank Coalfield. The No. 4 Seam in this area varies in depth of overburden from 75 m to 6 m, with an average of 55 m. Initially, the mine was planned as an underground mine incorporating a three-shaft system for transporting some 8 million tons of coal per annum to the power station. As the result of negotiations between Anglo and Escom to improve the volumetric extraction of coal from the new mine, the Kriel Division of Anglo Power Collieries was finally planned to produce 4,1 million tons of coal from an underground mine and 4,5 million tons from an opencast mine, with a resultant increase in volumetric extraction. This paper describes the opencast operation at Kriel, with particular reference to the water problems encountered and to the operation of the draglines on sloping ground to a depth of 40 m to the top of the coal seam.

Jan 1, 1980