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1. Book reviews Nonparametric geostatistics, by S. Henley. London, Applied Science Publishers, 1981. xiv + 145 pp. Reviewer: Douglas M. Hawkins Mineral deposits and global tectonic settings, by A. H. G. Mitchell and M. S. Garison. London, Academic Press, Geology Series, 198!. Reviewer: J. Lurie The mining and mineral laws of South Africa, by B. L. S. Franklin and M. Kaplan. Durban, Butterworths, 1982. Cloth, 240 X 150 mm, Lxxcii + 853 pp. R197,50 + GST. Reviewer: M. J. Martinson Aluminium- Volume II, by M. H. Govett and J. Larsen. Sydney, Australian Mineral Economics (Pty) Ltd, 1982. Reviewer: D. G. Maxwell International technology for the nonferrous smelting industry, by Noyes Data Corporation. Reviewer: D. A. Viljoen 2. New publications Coaltransport- an analysis of technical and economic factors, by Dag Pike. London, Financial Times, 1982. £86 (D.K.), D.S. $175 (overseas including first class airmail postage). Small solute concentration differences determined by interferometry, by H.E.L.G. Schweigart. Pretoria, CERG-CSIR (P.O. Box 395, Pretoria), 1982. 3. Mintek reports Report M8A determination, using solid zirconia electrolytes, of the activities of chromium oxide in ferrochromium alloys and slags at 1650°C. Report M31Software for the design of multivariable control systems by use of the Inverse Nyquist Array. Report M37A literature survey of the metallurgical aspects of minerals in Witwatersrand gold ores.This survey reviews the information in the literature Report M 39 The use of masking agents in the determination, by hydride generation and atomicabsorption spectrophotometry, of arsenic, antimony, selenium, tellurium, and bismuth in the presence of noble metals. Report M40 The use of X -ray-fluorescence spectrometry in the determination of uranium, cobalt, and sulphur in materials from mine dumps. Report M44 A n on-line dilution system for spectrometry ustng an inductively coupled plasma source. Report M47 The separation of trace elements from iron, and the analysis of iron-rock ores for use as reference materials. Report M49 A programme in BASIC for the appending of computer files. Australian coal The editors of World Coal and World Mining have announced plans to sponsor the IVth International Coal Exploration Symposium. The symposium will be held in Sydney (Australia) from 15th to 19th May, 1983. Small mine economics The editors of World Mining and World Coal have announced plans to sponsor the Second International Symposium on Small Mine Economics and Expansion in Helsinki, Finland, from 12th to 16th June, 1983. The first symposium was held in Taxco, Mexico, in May 1981.

Jan 1, 1982

By Y. J. F. Haven, G. H. J. Kitson, D. G. Beadle

I have read this paper with considerable interest. By using an automatic instrument to assess the dust samples taken in South African collieries, the authors are ahead of the gold mining industry, which still uses the laborious method of microscopic assessment, with its limited accuracy, of the half-a-million dust samples taken each year. As the authors state, the result obtained from the photoelectric assessment of the samples is approximately proportional to the surface area of the respirable dust. However, modern theory is that, in assessing coal dust, it is the respirable mass which is important in determining the danger to health. Thus Table II of their paper is of considerable value in showing the relationship between the P.E.R. value and the mass concentration measurement. Unfortunately this table only gives the mean relationship. The important point is to know the scatter around this mean. If the scatter is small, then this mean value (for different operations) would have considerable value in converting from respirable surface area to respirable mass, but if the scatter is large, then such a conversion, using a fixed factor, will have little value. In the text the authors mention the range of values as being between 3 to 24, but without a statement on the number of samples used to determine this range and their distribution, this is a rather nebulous statement. I hope, in their reply, that the authors will give, in statistical terms, some statement on the overall correlation between these two parameters. I might add that at present a member of my staff is working on similar relationships for dust samples taken in gold mines. Another factor which may well be most important in determining the danger of dust levels in coal mines is the composition of the dust, specifically its quartz content. I understand that the authors have some information on this point, and to make their results of greater value in International comparisons of dust levels and the concomitant development of pneumoconiosis, it will be extremely useful if they could place this information on record. Even though, as I have said above, modern theory postulates that respirable mass is a better parameter to use than respirable surface area, the problem remains of how to measure this parameter with short time sampling, which, as the authors point out, is their preferred strategy for dust sampling and dust control. I know of no easily portable instrument, independent of external power supplies, which can, in 5 to 10 minutes, collect sufficient dust to enable it to be weighed accurately. Thus the authors' implied intention of continuing to sample by their present techniques and using their present parameter to report the results is, in my opinion, an entirely valid decision.

Jan 10, 1968

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

THE BRITISH COUNCIL AWARDS 1971-1972 The British Council invites applications from permanent residents of the Republic of South Africa and South West Africa, for the following awards: SCHOLARSHIPS Scholarships are available to such permanent residents who hold university degrees or comparable professional qualifications. The scholarships are tenable for postgraduate study at universities and other educational establishments in Britain for the academic year of approximately ten months beginning 1st October, 1971. BURSARIES Bursaries are available to such permanent residents who hold university degrees or who, being already established in their professions or trades, hold other qualifications. The bursaries are tenable at British educational institutions or other suitable establishments in Britain for a course of training, study or research lasting not less than two months or more than six. The British Council may, at its discretion, bear the cost of single or return fares to Britain. The closing date for applications is 31st July, 1970. TRAVEL GRANTS Travel grants each of £300 are available in respect of journeys to Britain undertaken between 1st April, 1971 and 31st March, 1972 and are provided to facilitate visits to British universities by persons holding staff appointments or graduate research posts at South African universities. Visits must normally be of at least six months duration. The closing date for applications is 31st October, 1970. For further particulars and the relevant application forms please write stating the type of award for which you wish to apply to: The Cultural Attache, British Embassy, Pretoria. THE SECOND INTERNATIONAL CONFERENCE ON THE PROPERTIES OF LIQUID METALS Tokyo, September 1972 The Second International Conference on the Properties of Liquid Metals will be held under the auspices of The Japan Institute of Metals and The Physical Society of Japan from 3rd to 8th September, 1972 in Tokyo. The purpose of this conference is to continue and extend the efforts towards the objective of the First International Conference on the Properties of Liquid Metals at Brookhaven, 1966. The tentative Conference subjects presently planned are: Structure of Liquid Metals and Alloys; Atomic Transport Properties (Diffusion, Viscosity, etc.); Electronic States and Transport Properties; Thermodynamical Properties; Melting and Phase Transitions. Correspondence regarding the Conference and the request for the First Notice should be addressed to: Professor S. Takeuchi, Chairman of the Organizing Committee, The Research Institute for Iron, Steel and Other Metals, Tohoku University, Sendai, 980, Japan. PROCEEDINGS OF THE NINTH COMMONWEALTH MINING AND METALLURGICAL CONGRESS 1969 Volume 4 of the Proceedings of the Ninth Commonwealth Mining and Metallurgical Congress 1969- Physical and fabrication metallurgy-contains the 24 papers presented under this general heading at the Congress. Also included are discussion, authors' replies, name and subject indexes, and plenary addresses by Sir Val Duncan, O.B.E. and Sir Ronald L. Prain, O.B.E.

By Z-P. Song, Z-B. Jiang, A-N. Jiang

"In deep excavations under conditions of high ground stress, even hard rock undergoes creep and other time-dependent effects that may result in instability. This study involves an analysis of the typical damage to the surrounding rock in the deep tunnels of the Jinchuan nickel mine in China. In order to understand the creep mechanism, triaxial creep laboratory tests were conducted under high stress. The creep characteristic curves and the Nishihara model parameters of typical rock were obtained from these tests. The numerical simulations, with which the creep model is developed, were performed using Flac3D. The creep model parameters were back-analysed on the basis of field monitoring data. Finally, appropriate support schemes for deep tunnels are suggested. The creep deformation of rock increases with time and axial load. At the same load level, the creep deformation of unloading is greater than that of loading. The simulated deformations of the surrounding rock are coincident with those from in situ monitoring, indicating the applicability of the presented rock creep model in situations of high ground stress. The results constitute a reference for support design in deep mining tunnels.IntroductionThe excavation of underground caverns can alter the natural stress field of the surrounding rock. In deep engineering projects under conditions of high ground stress, even hard rock shows obvious creep properties and time effects (Yao, 2005; Liu, 1994). The time effect of excavation not only causes decompression of the floor and increases sidewall stress, but also leads to stress concentration at the corners of the excavation, which influences the stability of the surrounding rock in a very complex manner.Many researchers have studied the creep properties of rock. ‘Griggs (1939) showed that creep occurs in sandstone, shale, and silty sandstone rock under the load of 12.5%–80% of the failure load.’ Subsequently, much research has been published on rock creep tests, including the uniaxial compressive creep test, the triaxial compressive creep test, the discontinuity rock shearing test, and field tests (Wawersik, 1974; Amadei et al., 1980; Okubo et al., 1991; Jin et al.,1995). A rheological model is used to describe the complex rheological deformation of material. This model is based on various fundamental parameters including rock elasticity, plasticity, and viscous properties, which can be combined into a constitutive model to reflect the rheological properties of various types of rocks. Wang et al. (2014) studied the creep– damage–rupture characteristics of rock salt by applying a creep-damage model. Munson (1997) developed a multi-mechanism constitutive model of creep of polycrystalline rock salt based on steady-state creep and modified it to incorporate transient creep through work hardening and recovery. An elasto-viscoplastic approach was developed to simulate the timedependent nature of the fracture zone (Malan, 1999). Rheological models were proposed to understand the collapse of iron ore mines in Lorraine by Dragan et al. (2003)."

Jan 1, 2016

The statistical methods described in the previous chapter permit the estimation of the average value p of a deposit, and the calculation of confidence limits for this value. They are based on the assumption of independence of sample values. As soon as a reasonable number of samples has been obtained from a deposit, local valuation, or estimation of blocks of ore within the mineral deposit, becomes necessary. The assumption of independence of values must then be rejected. Local valuation is possible only if there is a relationship between the value of a sample and the value of the surrounding ore, provided this relationship is a function of the position of the sample with respect to the blocks of ore to be valued. Whichever method is used for the valuation of a block of ore, the following assumptions are always made: ? The values of samples located near or inside the block are related to the value of the block. ? The values of the samples located closest to the block are most closely related to the value of the block. These assumptions will hold true only if the following assumption can also be made: ? There exists a relationship between sample values which is a function of the distance between samples. We are saying that the value x (z) of a sample centred at a point z has properties which are functions of this point. A certain spatial structure exists in the sample distribution. We are therefore dealing with a regionalized phenomenon, and we must treat x (z) as a regionalized variable. This can be done statistically by means of a model chosen to represent the spatial structure of the phenomenon. We shall see that the relationship between block values and sample values can be deduced from the relationship between sample values, and all local valuation problems can be solved using the statistical model obtained from the analysis of the sample values. It is important to realize that this statistical model is obtained by analysing the properties of known sample values, and will be used for the estimation of unknown block values. Such a statistical inference is valid only if the model developed represents the properties of both the sampled section of the ore body and the section to be valued: some stationarity conditions must be satisfied, and the model with the least restrictive conditions is preferred. In the following chapters, the most important models used in spatial geostatistics are presented. The models which can be used when there is no significant drift (trends) in values are described in Chapters 3-1 1. These models can also be used to solve some specific valuation problems when a linear drift is present. A general study of the more difficult problem of analysis of an ore body whose values present a drift is described in Chapter 12. Note that the area considered for local valuation of a block of ore is often small enough for any drift which might be present to be ignored.

Jan 1, 1978

By A. J. Vietti

"Drought conditions and the possible lack of sufficient raw water supply are a constant reminder to the Southern African diamond miner of the need to minimize raw (new) water input while maintaining or expanding mine throughput.Since water input volume is directly related to water output (loss) volume, mainly to the tailings, it is obviously beneficial to dewater the tailings as much as possible. However, this may not always be as easily achieved as first thought. Owing to the unique mineralogy of each kimberlite deposit and the unique chemical profile of each raw water source, dewatering of the slurry is not a standard process that can be applied equally across the entire diamond industry.This paper presents a first-step strategy that kimberlite diamond mine operators should consider and apply to meet and exceed their watersaving targets. Introduction A common metric for measuring water use efficiency in a metallurgical plant is the amount of raw water used per ton of head feed treated (m3/t). The metric includes all water recycled from the dewatering process and from the tailings storage facility (TSF), and the only inputs are tons of ROM ore and cubic metres of raw water make-up to replace water lost to evaporation, seepage, and lock-up at the TSF (Figure 1).The volume of new raw water available to the process is not unlimited and a maximum monthly allowance is typically set during the mine permitting process. Consequently, if the volume of water lost exceeds the volume of new water available or alternatively, in times of drought where the allowable volume is unavailable, the plant has no other option but to reduce throughput.However, if the plant operator is able reduce the volume of water lost to the TSF, the plant will have additional capacity to either increase tonnage throughput or sustain nameplate capacity during periods of severe drought. The obvious conclusion, therefore, is to reduce the volume of water lost to the tailings by improving the efficiency of the dewatering process (typically gravity thickening) simply by increasing the density of the tailings being discharged. In practice, however, the unique mineralogy of each kimberlite deposit and the chemical profile of the raw water source used imply that at some mine sites, solid/liquid separation (or settling) of tailings slurries at the dewatering process is more easily achieved than at others. In these cases, simple practical steps such as firstly improving the flocculant reagent make-up and dosing control, and secondly focusing on thickener discharge control, are the only requirements for enhancing water recovery at the dewatering circuit. At other mine sites however, solid/liquid separation of the tailings may not be achieved at all, in which case efforts to increase water recovery by improving reagent dosage and thickener control would be fruitless unless the colloidal properties of the thickener feed slurry are modified."

Feb 1, 2019

By W. A. Hustrulid

CONTENTS PART I The theory of percussive rockdrill operation. Blow frequency Piston impact velocity and blow energy Thrust requirements Prediction of penetration rate Drillsteel life Bit wear PART II Measurements of percussive rockdrill performance Thrust-penetration rate measurements Blow frequency Piston impact velocity and blow energy Prediction of penetration rates The effects of hole diameter Thrusts applied by men Bit wear Appendix I Experimental procedure and rockdrill and drill steel specifications Appendix II Piston impact velocities and energies PART III Practical implications. Provisions of adequate thrust Conclusions INTRODUCTION In South African gold mines some 100 million tons of rock are broken each year by drilling and blasting. Only minor improvements to this method of rockbreaking have been made since the general introduction of pneumatic rockdrills for stoping 50 years ago and, more recently, of drillsteels tipped with tungsten carbide. The combined total cost of drilling and blasting amounts to almost R0,50 per ton of rock broken, which is just short of one quarter of the direct cost of stoping or development. Rockdrilling accounts for about R0,25 of the total cost of rockbreaking, of which RO,12 represents labour costs and R0,13 the cost of drillsteels, rockdrill spares and power. Mine air compressors use over 7 X 108 kWh of electric energy a year, amounting to 7 percent of the total electric power used by the gold mining industry. The efficiency of percussive rockdrilling is, therefore, of considerable significance to mining costs. Over the past four years the Mining Research Laboratory of the Chamber of Mines of S.A. has been engaged in experimental studies of percussive drilling. These experiments were conducted at test sites on Robinson Deep Gold Mine and Vlakfontein Gold Mine, made available by Gold Fields of South Africa, Limited. Three rockdrills, designated types A, Band C, of different manufacture, were used in the experiments which involved the measurement of the following operating characteristics: 1. Throttle air pressure 2. Applied thrust 3. Penetration rate 4. Air consumption 5. Blows per minute 6. Drill steel revolutions per minute. The results of these investigations are presented in three parts. Part I, "The theory of percussive rockdrill operation" which provides a concise, mathematical description of the performance of a percussive rockdrill in terms of its principal mechanical dimensions and operating conditions, and an analysis of the bit and gauge wear of drillsteels. Part II, "Measurements of percussive rockdrill performance" in which the experimental results are presented in terms of thrust-penetration rate curves at different throttle air pressures and throttle pressure-air consumption curves, which are shown to agree well with the predictions of theory. In addition, data relating gauge wear on the drillsteel and penetration rate to the distance drilled are discussed. Part III, "Practical implications" in which an analysis of the status of thrusting devices is made and the benefits that would derive from using uptimum thrusts are discussed.

By C. van der Merwe, R. Lumbule, J. Kgomo

"The mining industry in the Democratic Republic of Congo has seen a significant expansion in copper and cobalt production in recent years. The greater proportion of this expansion has arisen through processing of oxide ores. The design of metallurgical processes for these oxide ores is based on the conventional equipment and layout for processing of sulphide ores, and comprises milling followed by leaching, solvent extraction, and electrowinning. The physical characteristics of the oxide ores, however, provide the process design engineer with an opportunity to examine critically the principles on which equipment selection is based and to design processes better suited to these ores. In 2015, ERG Africa began a project to expand the leaching capacity of the Luita hydrometallurgical facility. A number of innovative changes to the conventional process design were included in the modified flow sheet, which were validated when the project reached design throughput in November 2017. Of these changes, the most novel presented an alternative approach to mitigate the impact of colloidal silica on the operation of the solvent extraction plant. Introduction ERG Africa operates the Luita hydrometallurgical facility located at Kakanda in the Democratic Republic of Congo (DRC), treating malachite and heterogenite minerals to recover copper as electrowon cathode and cobalt as a hydroxide precipitate. The process was originally conceived as a heap leach operation followed by direct electrowinning of copper, with solvent extraction (SX) and electrowinning (EW) of cobalt. In preparing the sized ore for the heap leach operations, the undersized ore fraction at a P80 of 625 µm was discharged to a fines storage facility for later processing. This fraction carried with it approximately 30% of the copper contained in the run-of-mine (ROM) ore feed, and an equivalent proportion of cobalt.In a subsequent development, the cobalt SX section was converted to a copper SX facility, the leached cobalt being recovered from a bleed stream from the copper heap leach circuit. The operating cost for the process as then installed placed this facility well in the fourth quartile of the industry cost curve. The initial Luita operation was established in 2010 at a time when metal prices were above long-term averages (copper above US$3.00 per pound; cobalt above US$20.00 per pound), and the drop in prices in 2015/2016 presented a major challenge. The London Metal Exchange (LME) copper price dropped below US$2.00 per pound, with cobalt dropping below US$10.00 per pound in Q1 2016. This set in motion a range of improvement initiatives for the Luita operation. Central to this was the installation of a tank leach circuit to process the discarded fines and a modified SX circuit to manage colloidal silica. The revised Luita circuit reached design capacity in November 2017 and has proven to be well suited to processing of the oxide ore feed. This process design challenges the boundaries of conventional thinking in the areas of leaching, solid/liquid separation, solution management, impurity management, and metal recovery, and is the subject of this paper."

Nov 1, 2018

By P. K. Van Manen

Anglo Platinum's Polokwane Smelter is situated outside Polokwane, in the Limpopo Province of South Africa. Wet concentrate is received from various concentrators along the Eastern Bushveld Complex. Approximately 60% of the total concentrate received is from UG2 Reef, and 40% is from Merensky Reef. The concentrate is dried in two flash dryers, and smelted in a single 168 MVA furnace. A PGM-rich nickel-copper matte is cast, crushed and transported to Rustenburg for converting. The plant was commissioned in March 2003. The wet concentrate, containing 12-18% moisture, is weighed, sampled by one of two auger samplers, and off-loaded into a concentrate shed, either directly through a grizzly into a hopper, or onto the floor. From the shed the concentrate is conveyed to 12 concrete silos. Each of the two identical flash dryers can be fed from six silos. From the flash dryers concentrate is transported pneumatically to a 3 000-ton dry concentrate storage silo. The flash dryers are rated at a nominal 66 tph of dry concentrate at 14% moisture, while the furnace can smelt 90 tph of dry concentrate. Therefore both flash dryers are required to operate when the furnace runs at design throughput. Each flash dryer consists of a hot gas generator (HGG), a disintegrator and drying column, and a section for gas/solids separation. In the HGG, gas heated up to 700° C is produced from the combustion of coal peas in a fluidized sand bed. The hot gas and wet concentrate are introduced to the disintegrator at the bottom of the drying column. The disintegrator breaks up any lumps of concentrate, and throws the concentrate up into the drying column, where it travels co-currently with the hot gas. A fan draws the hot gas from the drying column and through the gas/solids separation section, which consists of three primary cyclones, a multiclone and a high-temperature bag house. The dry concentrate reports to a 450-ton dry concentrate storage bin. A portion of dry concentrate is recycled to the wet concentrate feed in a paddle mixer, so that concentrate that is too wet does not enter the disintegrator. The drying column outlet temperature is controlled at 120° C. Coal is received by road, screened for oversize and conveyed to a 400-ton concrete silo. From here it is pneumatically conveyed to a small feed bin at each of the HGG' s. The optimization work consists of process control improvements relating to the HGG outlet temperature, to control the drying column temperature during the ramp-up of the HGG outlet temperature and bag filters. Furthermore a heat balance is used to evaluate options to increase the instantaneous drying capacity of the existing equipment. The coal consumption derived from the heat balance is compared with the actual coal consumption. The purpose of the optimization work is to achieve an operating schedule for the flash dryers such that they operate simultaneously for as short time as possible. Most of the time one flash dryer should operate, allowing maintenance to be done on the other flash dryer, while drawing down the dry concentrate stock. Keywords: Anglo Platinum, Polokwane Smelter, pyrometallurgy, metallurgy, platinum, platinum group metals (PGM), base metals, flash drying, fluidized bed.

Jan 1, 2006

By J. D. Stone, I. D. Brumby, J. D. Inch, C. J. Beukes

A. Continuous miners, and their operation at Bosjesspruit Colliery J. D. INCH*, B.Sc. (Min.), M. I. Min. E. (Visitor) A brief description is given of the main types of continuous miner and their incidence of application in South Africa. Definitions and descriptions of the design parameters and components of the main cutting head are given. together with an outline of the modifications carried out to improve machine cutting performance. An account is given of experience with the application of water infusion to suppress airborne dust during continuous-miner operations. B. The selection and performance of continuous miners at Matla Coal by J. D. STONE, B.Sc. Eng. (Min.,) M.B.L. (Graduate) The reasons for the choice of continuous miners for the development of accesses and the winning of coal in a modern underground coal mine are discussed. It is shown that, at relatively low depths, the extraction rate as well as the costs are more favourable than those of proved total extraction methods. The selection of a continuous miner is largely dictated by factors such as production rate, capital and running costs, imported content and the extent to which local content and manufacturing are to be increased, manpower requirements, seam thickness, expected floor and roof conditions, and maximum cutting height. The present inherent shortcomings of continuous miners, such as inefficient continuous transportation of coal away from the miner and the lack of an infinitely variable traction to match the sumping speed with the feasible penetration speed, are pointed out. Specific problems encountered and the methods of solving them, as well as a planned trial with a heavy-duty continuous miner, are discussed. C. Continuous mining within the Tavistock Group of collieries I. D. BRUMBY.Dip. Min. (Visitor) The Tavistock Group of collieries is situated some 25 km south of Witbank. The three collieries comprising the Group are Tavistock Colliery, Phoenix Colliery, and South Witbank Coal Mines. Continuous mining is practised to some degree at all three collieries. This system of mining was introduced at Phoenix during 1976, South Witbank during 1978, and Tavistock during 1979. For the twelve months ended June 1979, continuous mining accounted for 1,7 Mt of the 4,5 Mt hauled from underground. All the continuous miners introduced to date have been of the rotary-drum type. D. Pillar extraction at Usutu Collieries using continuous miners by c. J. BEUKES(Visitor) In 1969 Usutu Collieries began pillar extraction with conventional mechanized equipment. This was very successful, but certain disadvantages could be eliminated if continuous miners were used instead. In 1976 the first continuous miner was introduced for this purpose; this proved so successful that a second unit was introduced in 1977, and pillar extraction by conventional equipment was discontinued. By the end of 1978, 1,3 Mt had been mined from pillar extraction by continuous miners. This paper describes the method in some detail, and indicates that the operational costs are lower than for conventional bord-and-pillar mining. Defect solid state A one-day multidisciplinary symposium entitled 'The Defect Solid State '80' will be held under the auspices of the Solid State Physics and Materials Science Subcommittee of the South African Institute of Physics at the University of the Witwatersrand on Monday, 14th July, 1980, following the Twenty-fifth Annual Conference of the South African Institute of Physics.

Jan 1, 1980

By Stephen Yirenkyi

The mining industry remains the backbone of many economies in the developing world. Its resurgence in Ghana since 1989 was driven by the global paradigm which emphasizes private sector-led development as the engine of economic growth in developing countries. The historical importance of mining in the economic development of Ghana is evident in the country?s colonial name, Gold Coast (Akabzaa T. and Darimani A., 2001). Ghana earned $2.5 billion from mineral exports in 2007. 20,000 people are employed in large-scale mining whilst 500,000 are employed in the small-scale sector. Mining contributes about 7% of Ghana?s total corporate tax earnings, 41% of total exports, 12% of revenue collected by the Internal Revenue Service and 5% of Gross Domestic Product (Ghana Chamber of Mines, 2008). Gold dominates the mining sector and the country is Africa?s second largest gold producer after South Africa. Surface Mining, which refers to a method of extracting minerals from the earth by their removal from an open pit or borrows, requires a large tract of land for its operations (Wikipedia, 2008). There is therefore a continuous divergence of interest between land required for surface mining and other land uses such as farming and housing, among others. In recent years, surface mining has been promoted in many mining countries because of the following reasons: ?Cost considerations compared to underground mining; ?Safety considerations, compared to underground mining; ?Low grade ore which requires processing huge quantities; ?Location of the ore bodies; and ?Competition among gold producing countries for investors. This paper looks at the socio-economic impacts and challenges of surface mining by using Gold Fields Ghana Ltd. Tarkwa, Ghana as a case study. The company has a concession of about 208 sq km located in the heart of the Wassa communities in the Western Region of Ghana and undertakes surface mining like most of the other mining companies in the country. Gold Fields, Tarkwa mine produces 700,000 ounces of gold a year with the potential of increasing this output to 800,000 ounces after its current expansion programme. Some of the effects of surface mining identified were: 1. Relocation/Resettlement of host communities and its negative effects; 2. Disturbance to Flora and Fauna; 3. Disturbance to Sacred Places; 4. Land Degradation; 5. Noise and Air Pollution; and 6. Water Pollution. This paper sought to investigate the measures put in place by the government and the Gold Fields Ghana Ltd. to mitigate the above effects in order to ensure peaceful coexistence between the mining company and its host communities. It concludes that most mining companies presume the provision of socio-economic development in mining communities will lead to the company securing a social license to operate. This however is not true but rather effective community relations combined with socio-economic development. According to Gold Fields Ghana Ltds. model, this involves continuous community engagement, conflict resolution and comprehensive environmental monitoring programme which will all lead to effective community relations. The strategies to ensure socio-economic development involves community development programme, livelihood restoration programme and community training and employment. All these will then lead to sustainable development which will ensure the company?s social license to operate. Gold Fields? model is in line with the principles of responsible mining. With this model, the community feels they are part of the company and the social license of the company is guaranteed. Gold Fields? model, which has been tried and tested for the past five years, could serve as a template with suitable modification for local conditions for other mining countries where surface mines operate.

Jan 1, 2008

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 'opening' 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 R. J. Davidson, M. V. Nkosi, V. Veronsee

When activated carbon is used for the recovery of gold and silver from Witwatersrand gold-plant solutions, severe fouling of the carbon by calcium carbonate is experienced. Also, difficulty is encountered in the retention of the adsorbed silver on the carbon because it is displaced by gold. The acidification of plant pregnant solutions before they are contacted with carbon was therefore investigated as a means of overcoming these shortcomings. The results from batch laboratory experiments on the effect of pH on gold adsorption from plant solutions and from synthetic solutions of high ionic strength indicated reaction rate optima in the region of pH 5 and below pH I. The effect of pH on the equilibrium capacity constant for gold adsorption onto activated carbon was also studied, using both a plant solution and a synthetic solution of high ionic strength. In the case of the plant solution, the results suggested only a marginal change in carbon capacity for gold in the pH range 12 to 6, with significant increases in capacity at lower pH values. In the case of the synthetic solution of high ionic strength, the capacity constant was found to increase steadily as the pH was lowered from 12 to I. It was also found that the gold adsorption capacity of the carbon was significantly better for the synthetic solution. No precipitation of gold from the pregnant solution (gold content approximately 6 g/t) was noted as the pH was decreased to values as low as ,. However, about 25 per cent of the 0,6 g/t silver content was found to have precipitated when a pH value of 3,0 was attained. Further substantial precipitation of silver took place as the pH was lowered further. The investigation was extended by the passing of suitably acidified gold-plant solutions through one or more columns containing activated carbon. Preliminary results showed that the particle size of the carbon had a significant effect on the rate of gold adsorption. No displacement of adsorbed silver by gold was observed to take place at an influent pH of 5. In a typical continuous 'merry-go-round' operation involving three columns in which gold was eluted from the lead column every 48 hours, effluent values averaging 0,001 g/t gold could be consistently achieved under simulated plant-operating conditions. The results were obtained at relatively high velocities of influent solution, resulting in very little hold-up of gold in the carbon-recovery circuit. More than 99,9 per cent of the gold and more than 99,8 per cent of the silver was recovered in the adsorption circuit, while the elution data indicated an average elution recovery of 99,8 per cent of the adsorbed gold and 98,2 per cent of the adsorbed silver. Elution of the carbon was carried out at 91 QC. This involved pretreatment of the carbon with a sodium cyanide solution, followed by elution with deionized water. At a loading flowrate of 66 bed volumes per hour, the indicated take-up of metals on the carbon in the lead column was as follows; 53 k g/t gold, 6 k g/t silver, 57 k g/t copper, 14 k g/t nickel, and I k g/t zinc. Only a trace of cobalt was adsorbed. At an elution rate of 1,7 bed volumes per hour, the lead column was found to be completely stripped in under 5 hours, with concentrations of gold in the eluate reaching 6 to 7 g/l. Besides the excellent adsorption and elution results obtained when acidified pregnant solution was used, no fouling of the carbon was indicated, and the pressure drop across the three columns was considered to be well within practical limits. When compared with conventional zinc precipitation, the carbon process would appear to have considerable economic merit, with strong indications that significant savings would be possible both in capital and operating costs. Further pilot-plant experiments, preferably on a gold plant, would be required before a meaningful cost comparison could be made with the existing process.

Jan 1, 1979

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

By Karl-H. Daum

Over many decades, numerous feasibility studies have demonstrated, that the production of sulphuric acid remains the most viable option of sulphur recovery from smelter off gas and abatement of SO2 emissions to the atmosphere. This is particularly more pronounced as smaller but more concentrated off-gas flows are to be treated from smelters, and enhanced sulphuric acid processes be-come available. Off-gas handling systems represent a significant capital and operating cost burden to the metallurgical operation. Modern pyrometallurgical smelter processes for sulphide ores based on the use of oxygen-enriched air, produce relatively small off-gas flows with high SO2 concentrations in the smelter gas of 30?60 %-vol. of SO2. This is a prerequisite for substantial cost reductions in the smelter off-gas handling and treatment system. As an alternative to sulphuric acid production, numerous scrubbing concepts with alkali or dual-alkali combinations as well as organic absorbents have been proposed. Also the reduction of the SO2 to elemental sulphur has frequently been studied. Very few of those alternative processes have been built in industrial scale, but all were generally characterized by none sustainable operation due to cost reasons, problems with issues related to chemicals used and by-products or poor availability. Thus the traditional concept of converting the SO2 to sulphuric acid is most common, although regarded as un-economic, but is at least a proven, environmentally sustainable and reliable way of sulphur gas processing. With the high acid price levels these days, operating companies are even able to generate significant revenue with their otherwise ?fatal? acid. Even though smelter gas is available at high SO2 concentrations, in a conventional acid plant one would add large amounts of air to dilute the gas down to a suitable concentration of 12?13 (14) %-vol. SO2. Dealing with higher SO2 concentrations in smelter acid plants is not feasible, as the gas exit temperature of the catalytic oxidation step would exceed the allowable limit for the catalyst. Hence those conventional acid plants are characterised by the use of large equipment, i.e. high capital cost, high energy consumption and limited flexibility. The newly developed LUREC® process can handle off-gas with significantly higher gas concentrations, even in excess of 25 %-vol. SO2. It is entirely based on well-proven equipment and unit operations. The first industrial application, operating with 16?18 %-vol. SO2 will be presented in this paper, along with the fundamentals of the process. It will be demonstrated that the process requires inherently lower capital cost, fewer operating cost, offers better energy recovery and lower emissions as compared to conventional design. While most existing smelter acid plants would have some built-in spare capacity, any significant increase of smelter capacity, say of 30%, can basically not be accommodated without installing an additional parallel new acid plant unit. An add-on LUREC® module can remove any such restriction, while simultaneously de-bottlenecking the existing acid plant. The paper also discusses the present status of the add-on technology for smelter acid plant expansions with respect to technical boundary conditions of different process alternatives as well as economic and environmental aspects. Application of the LUREC® process to modern smelter acid plants operating with high SO2 gas concentration, will lead to a substantial reduction of the specific size of the gas processing equipment and equivalent savings in capital and operating cost. The LUREC® process is patented worldwide by OUTOTEC.

Jan 1, 2009

By Axel Schulze

HUGO PETERSEN has developed from the traditional firm of the same name that was established in 1906 in Berlin and is an engineering company specialising in general plant engineering now based in Wiesbaden. HUGO PETERSEN has many years wide-ranging experience in process plant engineering, primarily in the field of Sulphuric Acid Processes and Gas Cleaning. Together with its main shareholder, Chemieanlagenbau Chemnitz GmbH (CAC), a company with over 40 years experience in the international plant engineering business, HUGO PETERSEN implements turnkey new plants from initial consultancy right through to commissioning. In addition, HUGO PETERSEN can offer its clients improvements and revamps of existing plant systems. Initial consulting, planning and implementation take place within the context of a lean company organisation with experienced engineers and specialists. In all respects, many years of experience guarantee expertise over a wide range of possible applications For clients from the chemicals and metallurgical industries, as well as those in metals processing, HUGO PETERSEN GmbH?s breadth of experience means not just an efficient introduction to new methods and processes, but also their full implementation. In addition, along with design and construction for individual components, the company?s range of services also covers supply of complete plants, as well as modernisation projects and revamps in line with local conditions and existing quality requirements With own technologies and patents in the field of gas cleaning and sulphuric acid, supplemented by licences and cooperation projects, HUGO PETERSEN ensures expertise over a wide range of applications. With specialists and expert knowledge in these fields of application, HUGO PETERSEN is your partner of choice, supporting you and bringing your acid, and project to a successful conclusion.

Jan 1, 2009

By D. F. Malan, W. J. Theron

Practical mining aspects should be considered when conducting pillar designs for bord-and-pillar layouts. The current methodology for pillar design will result in increasing pillar sizes with depth. This affects the extraction ratio and will result in onerous ventilation requirements when cutting large pillars. A holistic approach, including all mining engineering requirements, is required to ensure that the rock engineering designs are optimized to ensure efficient mining operations and sustainable production. Bord widths should not only be a function of the rock mass ratings, but should also be selected to fit the specifications of the mechanized equipment. The use of a ‘squat pillar’ formula for hard rock is discussed in the paper and the formula based on the exponents of the Hedley and Grant pillar formula, is explored. The effect of abutments and geological losses on average pillar stress is also explored. These factors must be considered when designing layouts at increasing depths.

Nov 27, 2024

By A. L. Edwards

A retrospective study was conducted on the Rand Mutual Assurance (RMA) noise-induced hearing loss (NIHL) compensation claims from 1998 to 2008 to determine if the 2008 milestone agreed at the 2003 Mine Health and Safety Summit was achieved. The number and costs of NIHL compensation claims in different commodity sectors and workplaces were collated. A secondary analysis of the ages of employees compensated after 2008 was conducted. A complementary retrospective analysis of audiogram data investigated the percentage loss of hearing (PLH) shift in different homogeneously exposed groups and occupations at two gold mines. The compensation claims analysis indicated a significant decrease in NIHL claims from 1998 to 2008, but the milestone was not achieved. The reason may be either that claims have not been submitted timeously as required by Instruction 171 and that the current submissions are a result of pre-2003 noise exposure, or that employees who were baselined are still developing NIHL owing to ineffective hearing conservation programmes in place in the mining industry. On the basis of best practice for hearing conservation, recommendations are made for leading indicators in hearing conservation programmes and for reducing the risks of NIHL in order to achieve the 2013 milestone.

Jan 1, 2012

By I. Zh. Bunin, M. V. Ryazantseva, V. A. Chanturiya

The paper proposes a new method for selective calciferous separation. The effect was achieved through the application of the combination of high-power electromagnetic pulses (HPEMP) and a specific reagent regime of the flotation. The combined application of HPEMP with the special reagent regime allowed for the improvement of the scheelite floatability: the recovery of CaWO4 rises from 41.2 % (untreated sample) to 58.5 (sample treated at 103 pulses). In contrast, the floatability of calcite decreases significantly: the recovery of calcite drops from 74.7% (untreated sample) to 41.1 % (sample treated at 103 pulses). The conditions were found for the selective separation of the minerals with very close flotation properties. Keywords: Calcite, scheelite, flotation, high-power electromagnetic pulses, DRIFT, XPS

Jan 1, 2020