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

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

Jan 1, 1989

By A. Becker, R. Tait, S. L. George, A. Carlisle

Surface treatments like laser shock peening (LSP) induce residual stresses penetrating up to 1 mm deep, surpassing traditional methods, thus significantly increasing the potential impact on materials prevalent in the automotive and aerospace sectors. Through extensive fatigue testing and fractography investigations, it was determined that LSP resulted in a fatigue life increase of over five times for aluminium alloy AA7075-T6 and titanium Ti-6Al-4V. Increased fatigue life was also noted in AISI316 stainless steel, which showed increases in fatigue life of 1.7x. While fatigue life in all cases improved, the efficacy of LSP was shown to be material dependent.

Jul 23, 2025

By E. Wroth, D. G. Krige

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

Jan 5, 1968

By L. A. Vermeulen, C. P. Campbell, D. D. Howat

The production 01material liner than 75 µm in a number of open-circuit mills was investigated. The mils ranged from laboratory rod and ball mils treating a variety 01 ores to large industrial rod mils treating gold and copper ores. It was shown that the dependence on feed rate 01 the production 01 material liner than 751J.1T1 can be accurately described by a simple formula that was derived in previous work. In some 01 the systems reviewed, it proved possible to assess the feed rate 01 ore to the mil at which the rate 01 production 01 material finer than 75 µm can be maximized.

Jan 1, 1991

By N. B. Strydom, P. L. Jooste, H. F. Kotze, W. H. Van Der Walt

Note: A revised method of heat acclimatization by N. B. STRYDOM, W. H. VAN DER WALT, P. L. JOOSTE, and H. F. KOTZE INTRODUCTION Climatic-room acclimatization was introduced in the gold-mining industry in 1965 and was rapidly adopted by all mines because it reduced the number of days required for heat acclimatization from 12 to 8. The method has been adequately described in previous papers1, 2 but, briefly, requires four hours of heat exposure per day while the work rate is progessively increased from 35 W on Day 1 to 70 W on Days 7 and 8. Several attempts to shorten

Jan 6, 1976

By G. van Aswegen

The geometry and morphology of a set of low-angle fractures around a stope in a deep Witwatersrand gold mine are explained in terms of extension fractures forming under variable conditions of stress. Primary extension fractures (E1) form some distance ahead of an advancing stope along the s1, s2 plane. With stope advance, a couple of these fractures end up in a stress regime conducive to transpressional shear and a secondary set of extension fractures(E2) is formed at a high angle to the primary fractures. i.e. at a low angle to the stope. As the E2 fractures are undermined, they migrate into a stress regime of transtensional shear and a third set of extension fractures (E3) may develop between E2 fractures. These have sigmoidal shapes, being parallel to the E2 fractures at the E2 discontinuity where s3 is negative, and curved through the un-fractured rock between E2 fractures where s3 is positive at the instant of fracturing. The fractures all display fractographic features characteristic of dynamic extension failure with striae indicative of the direction of rupture propagation and the local, instantaneous orientation of s1.

Jan 1, 2012

By T. Nghipulile, G. Dzinomwa, J. Kurasha, M. Amwaama, B. Mapani, K. Maweja, T. E. Moongo

The effect of mineralogy on the grindability was investigated using three copper ores – two sulphides and one oxide. The dominant copper minerals were identified by optical microscopy and mineral chemistry derived from SEM-EDS analysis. The sample designated sulphide 1 was borniterich, sulphide 2 ore was mainly chalcopyrite, and the oxide ore was predominantly malachite and minor azurite. The gangue minerals were identified using semi-qualitative XRD analysis. Sulphide 1 contained more than 80% (w/w) of quartz compared to about 70% in the other two ores. The Bond work indices were 13.8, 21.6, and 17.3 kWh/t for sulphide 1, sulphide 2, and oxide ore respectively. This suggested that the chalcopyrite-rich ore is the hardest, while the malachite-rich ore has intermediate hardness, and the bornite-rich ore is the softest. The brittleness indices of the ores were calculated using the chemical composition of the gangue, and a good correlation between brittleness indices and Bond work indices was observed, which highlights the importance of the gangue composition in determining the fracture behaviour of the ores. There is scope for further investigation into the relationship between ore mineralogy and comminution behaviour using other breakage characterization techniques.

Mar 30, 2023

By M. B. C. Brandt

The near-surface wave attenuation factor k (kappa), which describes the attenuation of seismic waves with distance in the upper 1–3 km of the Earth, was determined for the Far West Rand gold mining area using seismic data recorded by station Parys (PRYS) of the South African National Seismograph Network. Twenty micro-events in the magnitude range 0.7 ? ML ? 1.8 for the period 1 July to 15 November 2015 were analysed. For the analysis a 10-second window for the S phase portion of the vertical component seismogram was selected. The result was an average k = 0.048 ± 0.014 s, which is much higher than for a stable continental region where k = 0.006 s. This is because the assumed Brune source model is inappropriate for describing mining-related events that often have moment tensors with a volumetric (implosive) component. The average k was calculated to be 0.098 ± 0.038 s for explosions of a similar magnitude. This higher value derived for mining-related k may be employed as a means of correcting the Brune source spectrum to calculate seismic moment as well as corner frequency for events in the mining areas of South Africa.

Jan 1, 2017

By V. N. Medvedev, A. M. Bryukhanov

Stationary mine air control equipment enables monitoring of CH4 concentration only in sensor locations. CMC-4 individual indicator built into a caplight has been developed in Ukraine, extending CH4 monitoring to manned job sites. Major technical characteristics of the indicator are demonstrated. It features two operation thresholds: 1st threshold 1.0 to 2.0% vol. 2nd threshold 1.5 to 2.5% vol. Basic absolute accuracy ± 0.2% Caplight battery is a common power supply for the lamp and CH4 indicator, providing at least 10 hours of the system non-stop operation. Average pre-failure life is at least 10,000 hrs. It features: an explosion proof design and intrinsically safe test circuit; a small-sized sensor as an integral part of the caplight; a hybrid microcircuit to facilitate functioning of methane monitoring unit. The CMC-4 indicator weight is the weight of the caplight itself and equals 2.4 kg. A two-threshold design makes possible to distinguish between methane sources: slow accumulation due to diffusion or sudden rushes due to a gas-dynamic event. CMC-4 indicators in addition to stationary equipment minimize the risk of CH4 explosion underground determined as a product of accident intensity by calculated time interval.

Jan 1, 2003

By S. Fletcher, C. Calle, K. Constanti-Carey, R. Ruzbacky, M. D. Horne

Laboratory test-work has established that two classes of redox catalysts based on anthraquinone and bipyridinium derivatives are effective in accelerating the rate of removal of metallic iron from reduced ilmenite (RI) in the aeration step of the Becher process. Two water-soluble anthraquinone derivatives are efficient: 9,10-anthraquinone-2,6-disulfonicacid, disodium salt (catalyst AQ-2,6) and 9,10-anthraquinone-2-sulfonic acid, monosodium salt(catalyst AQ-2). When 0.2 per cent w/v of either was added during aeration the rate of metalliciron removal was accelerated by two to three times. Furthermore, the addition of either compound resulted in magnetite (Fe3O4), the iron oxide phase preferred in commercial practice, being formed. However, both AQ-2,6 and AQ-2 were rapidly lost from solution during aeration although their catalytic activities persisted even when their concentrations in solution were low(<0.05 per cent w/v). The negatively charged anthraquinones were lost from solution most probably by adsorption onto the magnetite that has a net positive charge at the pH values encountered during aeration. To address this problem, the catalytic activities of positively charged bipryridinium derivatives were tested. Two water-soluble bipyridinium compounds, paraquat (1,1’-dimethyl-4,4’-bipyridinium dichloride) and diquat (6,7 dihydrodipyrido [1,2-a:2’,1’-c] pyrazinediiumdibromide), were found to significantly accelerate the rate of removal of metallic iron from reduced ilmenite, and both gave magnetite as the oxide product in the laboratory tests. Moreover, recycle tests showed they were not lost from solution as quickly as were the anthraquinone derivatives. Results in larger scale plant laboratory reactors also showed a significant increase in aerationrates when diquat was added and magnetite was again the oxide formed. However, the oxygensupply to the reactors was insufficient to achieve the maximum effect of the catalysts on theaeration rates. The limited supply of oxygen in a full scale aerator may prevent the application ofthe catalysts in a commercial Becher operation. Keywords: Becher process, aeration, anthraquinone, bipyridinium, redox catalyst, reduced ilmenite

Jan 1, 2003

By S. Subramanian, L. Mweene

A siliceous low-grade copper ore from Mpanda Mineral Field (MMF), located in south west Tanzania consisting of SiO2 (58.3%), Fe (10.4 %) and Cu (0.9 %) was initially subjected to selective dispersion-flocculation using sodium trisilicate (STS) as a dispersant and xanthan gum (XG) as a flocculant. Subsequently, the selectively flocculated fraction was floated using a collector combination of potassium ethyl and amyl xanthate (1:1) in the presence of poly-ethylene oxide (PEO) as a silica depressant, to get a copper concentrate. The mineralogical investigations confirmed that the MMF ore was mainly composed of quartz, chalcopyrite and pyrite. Adsorption studies indicated that the adsorption density of XG on MMF ore decreased with an increase in pH, while that of PEO was independent of pH. The adsorption isotherms of XG and PEO on MMF ore exhibited Langmuirian behaviour. The zeta potential studies of MMF ore indicated that the iso-electric point (IEP) was located at pH 2.6, presumably due to its high silica content. In the combined presence of XG and STS, the zeta potential of MMF ore became more negative, with an increase in the concentration of STS. However, in the presence of PEO, the zeta potential of MMF ore became less negative with an increase in the concentration of PEO beyond the IEP, without shifting the IEP. Selective dispersion-flocculation of MMF ore improved the flocculated chalcopyrite grade to 19.8 %, yielding a degree of flocculation and a maximum separation efficiency of 12.1 % and 85.7 % respectively, after three desliming stages in the presence of 30 ppm of STS and 30 ppm of XG. A copper concentrate of 29 % grade and 90 % recovery containing silica of 0.9 % grade and 1.7 % recovery was achieved by flotation of the selectively flocculated component, using PEO as a depressant for quartz. Keywords: Flotation kinetics, flotation machines, flotation bubbles, particle size

Jan 1, 2020

By S. T. L. Harrison, M. Smart, C. J. Edward

Biooxidation, a biohydrometallurgical process, utilizes a mixed microbial consortium to mediate the dissolution of refractory gold-bearing ores to enhance gold recovery during cyanidation. Recognition of water as a critical resource has prompted improved water management strategies within mining and minerals processing. Bioremediated wastewater from the cyanidation tailings provides a potential water source to supplement the water required for the biooxidation circuit, reducing freshwater extraction. Gold biomining operations have typically been reluctant to recycle this wastewater due to the potential inhibitory effects of cyanide and thiocyanate (SCN-) on the microbial population. Historical process upsets suggest that the microorganisms are sensitive to SCN- concentrations as low as 1 mg/L. However, effective bioremediation processes resulting in effluents with low or negligible SCN- concentrations may allow gold biomining plants the opportunity to recycle available water. Microbial activity as well as alteration in community structure were monitored as a function of increasing SCN- concentration over the range of 0 mg/L to 5 mg/L to investigate the response of the biooxidation community to potential SCN- spikes resulting from upsets in the remediation process. Appropriate microbial growth and ferrous iron oxidation activity with a nominal shift in community structure were observed at SCN- concentrations up to 1 mg/L. Thiocyanate concentrations above 1 mg/L resulted in the complete inhibition of ferrous iron oxidation activity. This was attributed to Leptospirillum ferriphilum’s sensitivity to SCN-. Exposure to a SCN- concentration of 5 mg/L resulted in a lower cell concentration and a significant increase in growth lag time, and the absence of L. ferriphilum from the microbial population. At. caldus and Ap. cupricumulans were still present at 5 mg/L SCN-, although Ap. cupricumulans was unable to oxidise ferrous iron at this SCN- concentration. It was concluded that recycling bioremediated cyanidation tailings wastewater within the biooxidation circuit is feasible; however, to retain process performance, SCN- concentrations should not exceed 1 mg/L. Keywords: Biooxidation, thiocyanate, water recycling, community structure

Jan 1, 2020

By F. Yang, S. S. Long

Mixed fatty acids, such as oxidised paraffin soap, are frequently used as a substitute for oleic acid in flotation of scheelite in China. Not only because it is cheaper than oleic acid, but also because it is superior in flotation performance to oleic acid. This superiority is generally attributed to the mixed fatty acids' synergistic effect. In this paper, the synergistic effect of oxidised paraffin soap was investigated. We found it mainly consists of a series of saturated fatty acid salts, with six to nineteen carbon atoms. All the contained fatty acids were classified into four groups based on their chain length to make the research work practical. Also, the groups were labelled as 1, 2, 3, and 4 in order. The research was simplified into studying the synergistic effect of four equivalent collectors, instead of fourteen fatty acids. It is justified by the fact that neighbouring fatty acids are expected to have a similar flotation performance. The synergistic effect of the four groups of fatty acids was studied by single-mineral flotation with single groups, binary mixtures, ternary mixtures, and the full component, as collector respectively. The flotation performance index and synergy value were introduced to evaluate flotation performance and synergistic effect of groups, respectively. The results show that the mixture 1+2+3 has the strongest synergy of all followed by the mixtures 1+3 and 2+3. The mixture 3+4 and 1+2, on the other hand, have negative synergy. The synergistic effect of mixtures of groups indicate that being consecutive in length and a broad range of lengths are vital for fatty acids to enhance synergy. Group 3 is fundamental for the collecting action of 733 oxidised paraffin soap in the flotation of scheelite. The fatty acid soap in group 4 is not suitable for flotation due to its low solubility. Keywords: Mixed collectors, synergistic effect, scheelite

Jan 1, 2020

By H. S. Baloyi, J. J. Rust

Spiral concentrators are widely applied to beneficiate commodities such as Cr2O3, mineral sands, and ferrochrome. The application of spirals for heavy minerals has been limited to a feed top size of 1 mm. At particle sizes larger than this, friction retards heavy particles and the particles' primary velocity approaches zero, resulting in a solids build-up, therefore inhibiting further separation. Recently, a new Multotec SC25 spiral model was introduced to the industry, which aids in beneficiating particles of up to 3 mm, with successes having been demonstrated in ferrochrome slag and titanium-bearing minerals. This spiral also has potential in electronic waste applications. This paper examines the applicability of the SC25 spiral in coarse Cr2O3 recovery. This stems from the fact that junior chrome miners tend to discard particles larger than 1 mm as they neither have milling nor dense media circuits. Also, alluvial chrome miners discard +1 mm even though this size fraction contains recoverable and liberated Cr2O3. The SC25 spiral has shown better Cr2O3 separation efficiencies than the SC21 model and with higher SiO2 rejection at saleable grades, thus providing an additional revenue stream specifically for the coarser (+820 μm) size fractions. Keywords: Spiral concentrators, chromite recovery, specific gravity, SC25 spiral, coarse spiral

Jan 1, 2020

By D. B. Swanepoel

Hulamin Rolled Products (Ltd) developed a high Mn and Si containing alloy, EN 4017 as a scrap consuming alloy from the waste generated in producing their range of clad products for brazing of inter alia automotive heat exchangers. The multifaceted composition of this alloy (Al -1.2 wt% Si -1.1 wt% Mn -0.2 wt% Mg-0.26 wt% Cu) suggested that it might display some degree of strengthening via precipitate formation of the Mg2 Si-types. The aging isotherms constructed showed that EN 4017 with 0.19% Mg reached a lower peak strength (YS~210 MPa and UTS~260 MPa)compared to the EN 4017 material with 0.43% Mg (YS~270 MPa and UTS~325 MPa). The higher Mg variant of EN 4017 was comparable with the age-hardenable reference alloy EN 6061. Selected area electron diffraction (SAED) studies proved that the microstructure of age hardened EN 4017 contained a combination of ß&apos;, ß" and U2 phase after being aged at 175ºC for 65 h. The general microstructure contained precipitate free zones as well as grain boundary Si films, although these did not lead to significant embrittlement. The role of Mn dispersoids in the fracture mechanism of 4017 is also discussed. The article will compare the behaviour of EN 4017 with that EN6061. Keywords: Age-hardening, dispersoids, magnesium silicide, precipitate free zones

Jan 1, 2009

ERRATA-AUTHOR&apos;S REPLY IN SEPTEMBER, 1970, JOURNAL: PAGE 52 I. Left half of the page, 4th paragraph, fifth line from the bottom: for 22 / 5 1/2 %, read 22 / 5 1/2 %. 2. Right half of the page, fifth paragraph from the bottom, third line from the bottom: for N(l - p), read N(1 - p)x. 3. Right half of the page, fourth paragraph from the bottom, second line: for: at time t after time zero, read: at time x after time zero. NOTICE TO ALL STUDENT MEMBERS OF THE SOUTH AFRICAN INSTITUTE OF MINING AND METALLURGY The Institute offers each year for competition among Student Members of the Institute a prize, or prizes, should entries so warrant it, of up to the value of R100 to be awarded to the Student Member or Members who submit the best paper or dissertation on a topic appropriate to the interests of the Institute. In the case of Student Members who are in full time study at a university the dissertation or thesis which they may have to write in part fulfilment of their university curriculum may be accepted provided it is presented in a manner and on a topic suitable for publication in the Journal. The closing date for the receipt of papers shall be 31st December each year.

By A. Modi, C. Li, R. Kabwit, W. Y. Xinchao, R. Joshi

The KIMIN copper project can treat 1.7 million dry t/a copper-cobalt ore. It has three mills in operation, two using conventional milling in water and the third in raffinate. We introduce the project process circuit and the capital and operating costs of using the raffinate process compared with conventional milling in water. There is a capital saving of approximately 11.5% for major equipment by removing the pre-leach thickener. In addition, it also achieves an annual operating cost saving of USD 2.1 million due to an increase in 49% of acidic raffinate recycled and reduction in lime consumption. The features of CITIC’s new duplex stainless-steel ball mill ,as the key equipment of the system, are also introduced

Jan 1, 2023

By K, J. MAUNDERS

This paper briefly describes a computerised model of the South African electricity system which has been developed jointly by the Atomic Energy Board and the Electricity Supply Commission. The results of an analysis of this system have been subjected to a fairly extensive sensitivity analysis which reveals that up to the year 1990 between 680 and 700 million tons of coal will be required for electric power generation, together with between 7 000 and 10 000 tons of uranium. This result is fairly insensitive to data variations, but the prediction of a cumulative requirement up to the year 2000, of between 1 300 and 1 800 million tons of coal and 25 000 to 70 000 tons of uranium, proves to be very data-sensitive.

Jan 1, 2014

By H. C. Voges

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

Jan 1, 1991

By E. J. Maunders

This paper describes briefly a computerised model of the South African electricity system which has been developed jointly by the Atomic Energy Board and the Electricity Supply Commission. The results of an analysis of this system have been subjected to a fairly extensive sensitivity analysis which reveals that up to the year 1990 between 680 and 700 million tons of coal will be required for electric power generation, together with between 7 000 and 10 000 tons of uranium. This result is fairly insensitive to data variations, but the prediction of a cumulative requirement up to the year 2000, of between 1 300 and 1 800 million tons of coal and 25000 to 70000 tons of uranium, proves to be very data-sensitive.