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By S. J. van Wyk, A. S. H. Haagner

Land rehabilitation transcends the mere establishment of vegetation across disturbed landscapes; it should be conceptualised as a specialised ecological reconstruction process. This process necessitates a long-term perspective on the success of land repurposing. It requires a continuum of inputs specified and managed by a dedicated and qualified custodian to ensure the sustained economic viability and ecological sustainability of the land. Based on the authors’ combined experience and field observations over more than five decades, they advocate for an urgent shift away from traditional land rehabilitation practices toward a paradigm centred on land capability reconstruction and active stewardship. The prevailing rehabilitation culture has overly simplified the process to a mere revegetation exercise, which is fundamentally incompatible with the comprehensive requirements for sustainable land capability. While the regulatory framework provides valuable guidance for obtaining a mine’s social license to operate, a paradigm reset is essential. Mining companies must prioritise rehabilitation quality, set achievable sustainability goals, and establish clear pathways and milestones to meet relinquishment criteria. Currently, the status quo view—a model reliant on outsourced contracting and agricultural approaches—is that mine rehabilitation is a one-time intervention rather than a complex, systemic, and long-term process. This perspective presents significant challenges, as the success of ecological interventions can only be accurately assessed over extended timeframes, often spanning decades. Shortsighted rehabilitation specifications, coupled with practitioners operating within agricultural decision frameworks, frequently overlook the necessity for integrated rehabilitation expertise and the scientific foundation essential for designing and implementing effective ecological outcomes. A lack of long-term monitoring information impedes the required adaptive management of rehabilitated systems and, therefore, delays the transition towards final land use. This short approach costs the mining industry millions annually, often without recognising the inadequacies in long-term liability mitigation and the substandard ecological success achieved. The disconnect between procurement processes, the qualifications of service providers, and the unclear pathways to relinquishment can impede successful rehabilitation outcomes, frequently undermining progress toward sustainability milestones. Unless there is a fundamental reset of the land rehabilitation culture, pursuing sustainability will remain an elusive and perhaps impossible objective. Consequently, mining companies may hesitate to fully commit to sustainable land rehabilitation practices, as the prospects for relinquishment and final closure remain uncertain at best.

Apr 25, 2025

By D. F. Malan, P. J. Le Roux

Almost no work has been published on the strength of pillars in mines exploiting the LG6/LG6A chromitite bands in the Western Bushveld Complex. The strength of these pillars is unknown, and the hard rock industry still uses the Hedley and Grant formula. Numerical modelling, using inelastic constitutive models, may be of some value in estimating the pillar strength, but this approach is difficult and prone to errors as many assumptions are made. This paper explores the alternative approach of the back-analysis of actual LG6/LG6A mining layouts, using displacement discontinuity codes to simulate the average pillar stresses on a mine-wide scale. From this, and underground observations, a ‘minimum’ pillar strength (K-value in the pillar strength formula) can be estimated. The codes TEXAN and Map3D are both used in the paper to simulate the actual pillar shapes and sizes in an LG6/LG6A mine. The results are encouraging as the two codes produced similar APS values for a complex pillar geometry. An estimated K-value of 77 MPa was determined for the LG6/LG6A pillars in this particular area. This is more or less similar to earlier work done for the UG2 chromitite seam, but higher than the K-value of 39 MPa for a linear formula obtained by other workers for LG6/LG6A pillars. Caution should therefore be exercised before using this higher value for all areas. The effect of element sizes and explicitly simulating the effect of the surface for shallow pillars is illustrated in the manuscript using both codes. The need for actual pillar observations and an iterative design process, which cycles between modelling and observations, is also emphasized.

Nov 27, 2024

By J. D. Swanepoel, S. Horn, C. J. Badenhorst

Diesel particulate matter (DPM) is the particulate component of diesel exhaust, which includes diesel soot and aerosols. When released into the atmosphere, DPM can take the form of individual particles or chain aggregates, with most in the invisible sub-micrometer range of 100 nanometres. The main particulate fraction of diesel exhaust consists of fine particles and because of their small size, inhaled particles may easily penetrate deep into the lungs. The rough surfaces of these particles make it easy for them to bind with other toxins in the environment, thus increasing the hazards of particle inhalation. DPM is a significant health hazard in underground mining operations, where the use of dieselpowered equipment in confined spaces can lead to high levels of exposure among people working in these environments. Classified as a carcinogen by the International Agency for Research on Cancer (IARC), DPM has been linked to severe health effects, including respiratory and cardiovascular diseases, as well as an increased risk of lung cancer. Effective management of DPM is essential not only for protecting people’s health but also for ensuring regulatory compliance, maintaining operational efficiency, and safeguarding the long-term viability of mining companies.

Mar 5, 2025

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

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

Jan 1, 2020

By J-N. Oyourou, R. McCrindle, C. J. S. Fourie

"Many mines in South Africa were inappropriately closed and left unattended during the last century. These old mines are potential sources of environmental pollution and may pose a health risk to local populations, since the surrounds contain elevated levels of toxic elements. Soil from the old Edendale lead mine property in Mamelodi Extension 11, east of Pretoria, which was suspected to be contaminated with lead, was analysed. The mine, which primarily extracted galena, was decommissioned in 1938. During this study, the old mine property and surrounds were screened for zinc and lead using a field-portable X-ray fluorescence (pXRF) spectrometer. The metal concentrations, which ranged from 18 to 7 300 mg/kg for zinc and from 50 to 21 000 mg/kg, for lead, were plotted together with the GPS coordinates to produce a concentration distribution map for lead and zinc. This map revealed three anomalies of unacceptably high concentrations, possibly corresponding to areas where ore had been piled or dumped to waste from the old mine. Thirty-six of the soil samples that were analysed by pXRF on site were selected and analysed in the laboratory using inductively coupled plasmaoptical emission spectrometry (ICP-OES) for verification purposes. The zinc and lead concentrations obtained from the two analytical methods were compared using linear regression analysis. Excellent correlation was obtained between the results from the two methods (r2 = 0.99 for zinc and r2 = 0.95 for lead). IntroductionEarly economic development in South Africa was based mainly on mining. However, these activities left a legacy of approximately 6 000 improperly closed mines scattered throughout the country (SAI, 2009; van Schie, 2012). Due to their proximity to settlements, these abandoned mines may pose a potential health risk to humans and an environmental hazard. Mining activities generate large amounts of waste, mostly in the form of overburden and process tailings, which often contain high concentrations of toxic elements. Bioavailable toxic elements can contaminate crops growing in the polluted area (Kala, 2010; Marwa, Meharg, and Rice, 2012) or eventually migrate and detrimentally affect the groundwater (Rösner and van Schalkwyk, 2000). Moreover, sulfides accumulated in mine waste generate acid minewater (Rösner and van Schalkwyk 2000). Unfortunately, owners of mines, closed before the promulgation of the Minerals Act more than twenty years ago (Act 50 of 1991), avoided the compulsory rehabilitation of the environment enforced by this legislation. Toxic elements, such as lead, which were confined within ores, may become exposed through mining and are therefore more easily mobilized within the environment. Elements essential to human and plant metabolism, such as zinc, when exposed by mining activities, may become hazardous due to elevated concentrations (Aslibekan and Moles, 2003; Wuana and Okiemen, 2011). Humans exposed to potentially toxic metals may experience harmful effects as discussed elsewhere (Nriagu and Pacynar, 1988; Wayne and Ming-Ho, 2005)."

Jan 1, 2019

The Institute has concluded an agreement with the Chamber of Mines whereby the Chamber Librarian maintains a section in the Library for the Institute. Surface effects of underground mining South Africa is blessed with vast mineral resources. It is essential to exploit those reserves if the country is to remain economically viable. Surface-mining equipment The Goldfields Mining Expo '86 is to be held in Kalgoorlie, Western Australia, from 30th October to 1st November, 1986, in conjunction with the Annual Conference of the Western Australian School of Mines. Mintek reports Report M142D Characteristics of gold and pyrite flotation at the Chemwes plant. (1st issued May 1984.) Report M159D Beneficiation tests on andalusite ore from the Marico district. (1st issued Aug. 1984.) Report M164D The recovery of tungsten from South African low-grade schee/ite concentrates. (1st issued Sep. 1984.)

Jan 1, 1986

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

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

Sep 9, 2024

[Consider the situation where we wish to estimate the value of a block of ore W from the values of n samples wt (i = 1, 2.... n). The samples can be located anywhere inside or outside W. The blocks and the samples can have any size or shape (Fig. 8.1). We define ws as the set of all the samples wi: WS = Iw1; W2; W3.. . wn and we wish to estimate the block W using the samples ws. Let xi be the value of the ith sample and µw the unknown average value of the block. We can estimate µw from the average value of all the samples. Let µw be the estimator of µw. As the samples need not be of equal size, the sample average is: µw = wi xi/~ wi. (8.1) The mean squared error made when estimating µw by µw is known as the variance of estimation of W by ws. It is the error made when giving the value of the sample set ws to the block W and is written o2E (WS to W). By definition: o2E (ws to W) = E [(µw - µw)2]. (8.2) If n = 1, this error is known as the extension variance. Since the sum of the weights given to the sample values is equal to 1,0, the estimator µw is unbiased (the question of the absence of bias is studied in more detail in Chapter 9): E [( µw - µw)1 = 0. (8.3) The error (µw - µw) has zero mean and a variance o2E = E (W8 to W). If we assume that this error is normally distributed, we can define confidence limits for µw. Let tP be the value of the standardized normal variance t such that the probability that t is less than tP is p: (t,) = .\/27T f exp (- t2/2) dt = p (8.4) The probability that µw is larger than µw + t1-P oE is p. The probability that µw is smaller than µw- t1-P oE is p. The following approximation formulae are often used to calculate 2p central confidence limits for the block value: 68% central confidence limits (p = 16%): µw - o2E, and µw + o2E. 95 % central confidence limits (p = 2,5 %): µw - 2 oE, and µw + 2 oE. Tables of the function 0 can be found in any textbook on statistics (for example, Fraser, 1958, p. 388) (see Table 7.2). [ ] The last line of Table 2.2, corresponding to f = 00, also contains the value of tP for p = 80, 90, 95 and 97,5 %.]

Jan 1, 1978

By W. Bleloch

Published in the Journal, September 1970, and presented at a colloquium on 'Future of the Steel Industry', 17th March, 1971. Errata (a) The final sentence in the penultimate paragraph, left hand column p. 34 of the text should read: 'In the hot regions of the bosh and stack the C+CO2=2CO reaction proceeds to CO from coke carbon and carbon dioxide generated by reduction of iron oxides by CO (the so-called solution loss of carbon)'. The italicised words have been omitted in the text. (b) Temperature in Fig. 3 is K not C. (c) In Fig. 9, connection from top plate of column 10 to heat exchanger 9 has been omitted. DISCUSSION J. B. BEEDON AND C. J. JONKER* Dr Bleloch's paper has outlined the advantages of recycling stack-gas of low nitrogen content in the blast furnace. In the first instance it would appear to be extremely beneficial in controlling the hearth temperature and allowing a 100 per cent oxygen-blast to be used. On a second point however we have to disagree. It is claimed that heat is liberated in the combustion zones by the oxidation of CO to CO2 near the tuyeres and in the subsequent reduction of the CO2 by CO. The author has hypothesized as follows, and we quote: 'The C+CO2=2 CO reaction in the outer combustion zone at about 1 700°C is no longer endothermic due to loss of stability of CO2 and its consequent dissociation into CO and free oxygen.' This statement requires closer examination. If we consider the individual events occurring with the CO in the recycled topgas, with the oxygen in the blast and with the coke in the burden, as done by Dr Bleloch, the following heat considerations emerge. All heats of reaction are calculated from the enthalpy values of each reactant and product, as well as the temperature dependence of their molar heat capacities. (a) At the tuyeres: CO+1/2 O2-+CO2 . . . . . (1) CO, CO2 at 500°C; O2 at 25°C. ?H reaction = -65 000 cals/mole CO The reaction (1) is exothermic. (b) In the outer combustion zone: CO2+C?2 CO . . . . . . . . (2) C, CO at 1 700°C; CO2 at 500 0C. ?H reaction = -56 000 cals/mole CO Reaction (2) is endothermic. The overall heat gain of reactions (1) and (2) is ?H=+56 000 - 65 000= -9000 cals/mole CO i.e. the overall reaction is slightly exothermic, the amount of heat liberated is however smaller than suggested by Dr Bleloch. (c) The dissociation CO2+CO+1/2 O2 . . . . . (3) is a highly endothermic reaction at 1 700 0C, requiring 82 000 cals/mole CO2 to proceed to completion. Furthermore, this reaction (3) is energetically not favoured to proceed, as can be observed from the free energy-temperature relationship of figure (3), which favours the formation of CO2 at 1 700 0C. Hence direct dissociation of CO2 at 1 700 0C appears to be unlikely.* The concept of recycling topgas is however still intriguing from many other points of view. Apart from providing a very efficient method of hearth temperature control, it will create a much higher CO partial pressure in the stack. Whether the rate of indirect reduction in the stack will be proportional to this partial pressure or not is un-

Jan 9, 1970

By P. J. D. Lloyd

The Federation of Societies of Professional Engineers (FSPE) recently conducted a survey of professional engineering manpower in South Africa. This article presents the data relevant to mining and metallurgical engineering. FSPE in their survey attempted to determine the supply of engineers in South Africa both from local universities and by immigration. The immigrants gave some problems in that it is difficult to know whether a man who claims to be professionally qualified really is. However, it was estimated that the ratio of immigrants who claimed to be qualified to those who really are the equivalent of our local graduates was between 1,0 and 1,9 for mining engineers, and between 2,0 and 4,1 for metallurgical engineers.

Jan 1, 1980

By A. W. S. Schumann

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

1. Reviews Steel and its heat treatment, by K-E Thelning. 2nd Edition. London, Buterworths, 1984.678 pp. R201,60 (incl. OS1). Reviewer: J.P. Hofman The Southern Queensland Conference 1985. Parkvile (Australia), The Australasian Institute of Mining and Metallurgy, 1986. A$ 15.00. Reviewer: R.B. McGilivray Rock mechanics, by A.R. Jumikis. 2nd edition. Clausthal-Zelerfeld (West Germany), Trans Tech Publications, 1983. US$ 58. Reviewer: R.W.O. Kersten The XVIth International Coloquium on Refractories. Freiburg, West Germany, 1985. DM 80. Reviewer: H.J. S. Kriek Diamond driling handbook, byW.F. Heinz. Johannesburg, South African Driling Association, 1985. 517 pp. R30. Reviewer: S.A.G. Martins Processing and utilization of high sulfur coals, edited by Y.A. Atia. Amsterdam, Elsevier, 1985.788 pp. Oft. 385. Reviewer: L.M. Falcon Weighingand proportioning of bulk solids, by Hendrik Colijn. 2nd edition. Rockport (U.S.A.), Trans Tech Publications, 1983. 398 pp. $60.00. Reviewer: G. Cunningham Investment appraisal and economic evaluation of mining enterprise, by S. von Wahl. Clausthal-Zelerfield (West Germany), Trans Tech Publications, 1986. US$ 40. Flotation of sulphide minerals, edited by K.S. Eric Forssberg. NewYork, Elsevier Science Publishers, 1985. 480 pp. US$1O1.75. Reviewer: T.R. Twidle 5th Meeting of the European Clay Groups (5th Euroc/ay Meeting), Prague, edited by J. Konta. Praha (Prague), Univerzita Karlova, 1985. Reviewer: D. Bibmann 2. Recent publications CSIRO. Research review1982-84: CS/Ra Division of Mineral Engineering. Melbourne, Commonwealth Scientific and Industrial Research Organization, 1985. 77 pp. The world copper market: structure and econometric model, by Gerhard Wagenhals. Berlin, Springer- Verlag, 1984. 202 pp. DM 37,00.Materials data of high-strength aluminium aloysfor durability evaluation of structures, edited by L. Schwarmann. Disseldorf (West Germany), Aluminium- Verlag GmbH, 1985. BRITISH GEOLOGICAL SURVEY. UnitedKingdom mineral statistics 1985, compiled by H.c. Squirel, A. Jones, N.E. Sharp, M. Lowe, and D. Slater. Keyworth (U.K.), the Survey, 1986. 175 pp.3. Mintek reports Report Ml66D The effect of calcining on the recovery of gold in a sand residue. (1st issued Sep. 1984.) Report M168D The treatment of Uitkomst concentrate. (1st issued Oct. 1984; reissued Mar. 1986.) Report M175D Supergene alteration at the Pering zinc-lead deposit: The nature and extent of the oxidation zone. (1st issued Oct. 1984.) Report M176D Bench-scale tests on the extraction of cobalt and nickel from COMAT leach liquors. (1st issued Nov. 1984; reissued Mar. 1986.)Report M241 PLA TO at Mintek. Report M243 The performance and wear characteristics of grinding media as affected by metallurgical and dimensional factors. Project development The Second Project Development Symposium on Planning, Construction and Commissioning will be held in Sydney, Australia, from 7th to 10th October, 1986. Hosted by the Sydney Branch of The Australasian Institute

Jan 1, 1986

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 A. Singh, R. R. Maclachlan

SYNOPSIS A back-analysis procedure based upon 35 mm photography was developed on a laboratory scale for the pur¬pose of determining the percentage of oversize particles known to be present in heaps of fragmented rock. Eight simple blends of crushed stone containing varying percentages of oversize on a 1:40 modelled scale were examined. Simulated test conditions for the following were subjected to the photographic study: (1) a bench heap, (2) the surface exposure of a partially excavated heap, (3) a gravity-sorted surface as occurs on an apron dump, and (4) four quadrant views of an incrementally built conical heap of rock fragments. The best back-analysis determinations were obtained from (4), the cone, over all eight blends of fragments. The mean error was within -0,16 per cent of the true value, with a range of ± 1,872. This was for oversize contents of 5 to 20 per cent where a minimum 25 per cent of the photograph area was processed by digitizer. The worst results were from the apron dump at - 1,462 per cent for a mean error of ± 4,174 per cent. The intermediate results were from the undisturbed bench at -0,186 per cent with a mean error of ±2,862 per cent.

Jan 1, 1989

By N. J. Alderman, D. J. Clowes, N. I. Heywood

The effects were studied of the presence of comminuted flint on the rheological characteristics of chalk slurry pumped through a 92-km long pipeline from Kensworth Quarry to Rugby Cement Works in the UK, and their impact on the pipe flow conditions. Flow curves were measured for 63different chalk slurry samples; these were described by the Bingham plastic model. In addition to observing trends in the yield stress parameter and plastic viscosity in the model with variations in both flint concentration and dispersant additive, use was also made of slurry viscosity values at a reference shear rate of 60 s-1for the pipe flow. It was found that chalk slurry viscosity decreases with increasing flint content up to 10% by weight for a given additive concentration, and increases with increasing additive concentration from 0.7 to 1.0%, for a given flint content. The pipeline flow regimes were identified using the critical value of the Bingham Reynolds number for laminar flow breakdown. For Bench 11 (quarry bottom) chalk slurries, it was observed that all sections of the pipeline operate in the turbulent flow regime for the nine different formulations and for the three pipe diameters used along the pipeline. However, this was not always the case for Bench 1(quarry top) slurries. It was therefore recommended that chalk slurries prepared from Bench 1 should not be used because of the strong likelihood that laminar flow conditions will occur owing to the presence of higher clay concentrations within this bench. This, in turn, will lead to an unsafe operation of the pipeline because of the potential for coarse particle settling. Likewise, blending of chalk from Bench 1 with the lower benches such as Bench 11 could be problematic and this requires further experimentation.

Jan 1, 2007

By H. M. Krieg, M. J. Ungerer, C. G. C. E. van Sittert, D. J. van der Westhuizen

"The transition metals tantalum (Ta) and niobium (Nb) are of significant importance, for example in the nuclear energy sector where they are used as cladding materials, as well as in capacitors and specialized materials. For these applications a high-purity metal is needed. The separation of Ta and Nb is always a challenge since they are found together in nature and have similar chemical and physical properties, resulting in costly and laborious separation processes. A technology that has been used successfully for the separation of these metals entails solvent extraction (SX)1. While separation was achieved in a previous SX study using a sulphuric acid (H2SO4) medium with the extractants diiso-octyl phosphinic acid (PA) and di-(2-ethylhexyl) phosphoric acid (D2EHPA), due to the absence of speciation data for Ta and Nb it is not clear how the separation occurred.One method that might be suitable for determining the speciation of a reaction is molecular modelling. Calculations based on the densityfunctional theory (DFT) are now used not only for light elements and small molecules, but also metal complexes, heavy metals, and especially metal separation in SX2. In this study the aqueous phase used during SX was investigated by studying periodic systems of Ta, as a metal and in salt form, when it is in contact with H2O and H2SO4. The results were used to predict the reaction mechanism occurring during SX. Results showed that (i) in a 1:1 acid-water ratio, the deprotonation of H2SO4 was endothermic, (ii) in a 1:5 ratio deprotonation was exothermic forming HSO4 -, and (iii) in a 1:10 ratio double deprotonation occurred to form SO42- exothermically."

Jan 1, 2017

By K. A. Annan

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

Jan 1, 2013

By J. C. Steinhobel

Answers to questions raised at the Symposium on Trackless Mining in Metal Mines Questions by J. O. Steinhobel (1) When rubber-tyred equipment is used for the delivery of stores and items of equipment, how is heavy material-pipes, rails, timber support, ventilation columns, etc.-dealt with? Termination of Membership In accordance with By-Law 7.6.1. of the Constitution of the Institute, the names of the undermentioned members have been removed from the membership register.

Jan 9, 1973

Colloquium and General Meeting A General Meeting and Colloquium was held on 'New Advances in Non-ferrous Pyrometallurgy' in Kelvin House, Johannesburg, on Wednesday, 17th March, 1976. The President, Dr R E. Robinson, opened the proceedings at 1OhOO,by welcoming the authors and delegates, who numbered approximately one hundred and twenty. Tinplate Conference The Tin Research Institute is organizing the 1st International Tinplate Conference, which is to be held in London from 5th to 8th October, 1976.

Jan 6, 1976

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

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