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By Sabine Verryn, Johan de Villiers

"Through the application of the Rietveld method, X-ray diffraction analysis has developed into a fully quantitative method for phase analysis. The Rietveld method itself has progressed from an unstable method that required constant operator intervention to a stable and robust method suited even for routine analysis. The method however, requires a good initial calibration and periodic checks on its precision and accuracy.Several examples from minerals processing, pyrometallurgy, and materials science show the strengths and applicability of the method. The limitations and sources of error are also discussed.1. IntroductionDiffraction of X-rays by matter has been used for more than 80 years, initially to elucidate the structure of crystalline matter, and to investigate the bonding between atoms. The first development of powder diffraction cameras and the later development of X-ray diffractometers saw the development of methods to identify phases and to try to quantify them in mixtures. A comprehensive database and search-match techniques have been developed over the ensuing years. Effective and accurate quantification had to wait for the so-called Rietveld' analysis where the total diffraction pattern is used and where the crystal structure data for the constituent phases are needed as input. Overlapping of diffraction peaks of different phases is handled without difficulty and reliable results can be obtained in spite of this occurrence.The development of methods for structure analysis from powder data has also progressed rapidly and more and more complex structures are being solved from first principles, i.e. without knowledge of unit cell or space group data. Because of the ease of use of the Rietveld method, it is finding increased application in metallurgy, especially in the quantification of phases of importance in minerals processing, in pyrometallurgical processes, and in materials characterization. This study will examine the principles of the Rietveld method, its application in metallurgy, and finally also the common pitfalls that need to be avoided.2. Rietve1d AnalysisAs the Rietveld method compares the experimental diffraction pattern with one calculated from known crystal structure data, the following data are needed:•,An XRD Pattern with as Iowa background as possible. This should have good resolution and high diffraction peak intensities. With modem X-ray detectors, this can be achieved in minutes rather than hours.•,Instrumental parameters such as goniometer radii and slit sizes are needed. These are the so-called ""Fundamental parameters"". They are used to calculate the shapes of the resultant diffraction peak profiles that are caused by the instrument. The parameters are used to correctly separate the instrumental effects on the diffraction pattern from the sample effects."

Jan 1, 2014

The following papers were presented at the colloquium* on `Mining methods and economics for improved coal extraction, and coal processing and preparation', which was held in Witbank on 3rd and 4th October, 1979. Review of Mining Methods for Improved Extraction of Coal, by P. King, of Gold Fields of South Africa, for¬merly of the Chamber of Mines of South Africa. The Application of Continuous Miners in South African Coal Mines, by J. D. Inch, Bosjesspruit Colliery, I. Brumby, South Witbank Coal Mine Ltd, J. D. Stone, Matla Coal Ltd, and C. J. Beukes, Usutu Colliers Ltd. Longwall Mining Experience at Coalbrook Collieries, by P. G. Henderson, General Mining and Finance Corporation Limited (with a contribution by C. J. Cloete, Sigma Colliery, Sasol I). Pillar Design in Coal Mines, by Dr H. Wagner, Chamber of Mines Research Laboratories. Open-cast Coal Mining at Kriel Colliery, by I. Buchan, L. F. Baars, and C. S. Northcote, Anglo Power Colliery, Kriel Division (with a contribution by M. F. Pleming, General Mining Coal Division). Richards Bay Coal Terminal, by B. Dunne, Richards Bay Coal Terminal. Coal Preparation in South Africa - a General Review,

Jan 1, 1979

1. Book review The geology of some northern Nigerian anorogenic ring complexes, by J.N. Bennet, D.C. Turner, E.C. Ike, and P. Bowden. The British Geological Survey: Overseas Geology and Mineral Resources, no. 61. London, HMSO, 1984. 66 pp. Reviewer: R.H. Ingram 2. Mintek reports Report M226 The separation of tungsten and molybdenum by the formation of sulphide complexes and extraction into a weak-base resin. Report M228 The solvent extraction of zinc, iron, and indium from chloride solutions by neutral organophosphorus compounds. Report M230 Shear flocculation of fines for improved flotation. Report M239 A mineralogical study of residue from the column leaching of a low-grade nickel-copper sulphide ore.3. Recent publication Driger Review, special edition, act. 1985. 48 pp. (Copies obtainable from Roly Nyman, HAMPO PMIE Division, P.O. Box 170, Joham1esburg 2000.) Black Mountain's achievement The 13 689 t consignment of metal concentrate that was loaded aboard the bulk carrier 'George' and shipped out of Saldanha Bay early in December brought the number of tons exported from Black Mountain via the Sishen-Saldanha railway and harbour system to one million.

Jan 1, 1986

By P. van der Westhuyzen

Gone are the days of stock-standard diamond beneficiation and final recovery circuits that could, with minor modifications, be adapted to treat any diamond-bearing ore source. A new era has dawned, offering the opportunity to streamline existing diamond processing operations or develop simpler, more efficient, and economical diamond processing plants, with the focus on more efficient comminution, beneficiation, and final recovery. This change in scene has been brought about by: ? The introduction, rapid development, and maturation of multiple comminution, sorting, and recovery technologies ? The need to adapt to a new standard of project approach post the commodity super-cycle phase, where optimizing existing operations and developing scalable, ?fit for purpose? new mines are fast becoming the norm in diamond processing plants in both primary and alluvial operations ? The quest for energy efficiency and lower labour costs ? More remote and inaccessible reserves (under lakes, tops of mountains etc.). This paper serves to identify some technology advances and demonstrates how these could be considered as replacements for or in combination with conventional technologies to arrive at an optimum techno-economic solution. To name a few applications/technologies: ? Comminution: conventional cone crusher, modified/specialized cone crusher, and the high-pressure grinding roll (HPGR). Although significant advances have been made in recent years, this paper only briefly covers comminution within the beneficiation circuit design ? Waste sorting: NIR (near-infrared) sorting, optical (colour) sorting, XRF (X-ray fluorescence) ? Primary concentration by combining dense media separation (DMS) with either XRT (X-ray transmissive), pulsed X-ray, jigs, or pan plants depending on the application, scale, and economics ? Final recovery of diamonds with either conventional X-ray technology, pulsed X-ray technology, or XRT. From recent studies, it can be concluded that there is no longer a standard solution, but rather the ?right? or appropriate solution. Through combining a sound knowledge of the ore source (ore dressing studies or onmine data gathering) and leveraging off the advances in technologies, one is able, through trade-off studies, to arrive at the ultimate techno-economic configuration, the ?right? solution. Emphasis is placed on maximizing diamond recoveries through appropriate technology selection and minimizing the associated costs in an effort to de-bottleneck or improve efficiencies of existing diamond processing plants, or to arrive at the ideal new diamond process plant design..

Jan 1, 2014

By K. -I. Higuchi, M. Mizutani, K. Haga, A. Shibayama, A. Batnasan

In recent years, the quality of iron ores has decreased due to the increase in the content of impurities, namely quartz, aluminum, phosphorus, and sulfur. It indicates that the removal of impurities from iron ore is an extremely challenging task. In particular, the high amount of phosphorus in iron ore could be a problem for energy consumption and CO2 emission in the steelmaking process. To reduce the impact of phosphorus, mainly two approaches: the utilisation of low-phosphorus iron ore, or dephosphorisation of high-phosphorus iron ore using mineral processing technology, have been suggested. In this paper, the removal of phosphorus from a low-grade high-phosphorus iron ore containing 0.1 mass% P and 57.5% Fe by alkaline roasting and wet grinding is discussed. The iron ore consists mainly of hematite, goethite, and quartz. Effects of sodium hydroxide concentration used as an alkaline source, roasting temperature, grinding time, and particle size on the dephosphorisation of highphosphorus iron ore were investigated. The dephosphorisation efficiency, iron loss, and mineralogical components of low-grade high-phosphorus iron ore under various experimental conditions were characterised by using XRD, SEM-EDS, and particle size distribution analysis. Results showed that the phosphorus content was lowered from 0.1 mass% to 0.034 mass%, while the iron grade reached 59.8 mass% from 57.5 mass% by alkaline roasting with 0.05 w/w NaOH at 300 °C for 2 h, followed by wet grinding in an attritor with water for 30 seconds at room temperature. Under the experimental conditions, the dephosphorisation rate was about 54.7%, and the solution obtained by wet grinding contained 22.7 mg/L of phosphorus and 21.5 mg/L of iron. The recovery rate of iron by this process reached 99.9%. Keywords: Iron ore, phosphorus, dephosphoriszation, alkaline roasting, wet grinding

Jan 1, 2020

By Hannweg L. A.

In a first for the diamond industry, Koffiefontein Mine in the Free State, South Africa, opted for the front cave mining method to extract a 90 metre high block of kimberlite ore. Block caving, because of its high initial capital requirements, resulting in a negative internal rate of return (IRR),was ruled out as a feasible mining method for Koffiefontein Mine. Front caving is essentially a combination of the traditional block cave and sub-level cave mining methods, but uses two or more levels of retreating drawpoints. Two extraction levels, in the case of Koffiefontein Mine, are spaced 12 m apart vertically, with the upper 478 m level (48Level) acting as the undercut/extraction level, and the lower 490 m level (49 Level) as the extraction level. Caving is achieved by undercutting a small column of ore, which is then drawn, on both the 48 and 49 Levels, until the planned extraction percentage is achieved. The next column of ore is then mined by advancing the undercut (48 Level), and the process is repeated until the entire ore block is depleted to the contact. Caving-by-gravity techniques of mining rely on closely controlled draw of the caved ore to ensure proper caving action and to minimize the dilution of ore with waste material from the open pit. This paper discusses the need for, and implementation of, a draw control system at Koffiefontein Mine — a computerized production management system, and a draw control system based on linear programming, to allow control over the draw rate and mining sequence — a major factor in all production planning to optimize production efficiency, ore recovery and cave management. Keywords: Draw control, Koffiefontein Mine, Front cave, Caving

Sep 1, 2001

By J. H. Fourie

This paper highlights modern practices in ore-treatment plants, with particular reference to the flotation of P2O5 at the plants of the Phosphate Development Corporation (Foskor) at Phalaborwa in the northern Transvaal. The following aspects are covered: design criteria, equipment and personnel, production standards and control. cost control, maintenance, and process development.

Jan 1, 1981

By Wm. Bleloch

SYNOPSIS The feasibility of rejection of nitrogen from stack gas by recycling of top-gas with cold-blast 95 to 98 per cent oxygen is assessed. It is concluded that nitrogen could be removed to a level below 3 per cent of the stack-gas using recycled cleaned cold top-gas and 95 to 98 per cent oxygen produced by a single-column air separation plant with liquid-air reflux instead of liquid nigroten reflux. The partial pressure of carbon monoxide in the stack-gas could be doubled without top pressure and the deposition of carbon and reduction of iron oxides by the stack gas could be markedly increased. It is concluded that rate of stock descent could be doubled compared with present-day practice and an iron make of 4400 kg/m 2 of hearth area/hour may be attainable with a working volume of 129 m2/m hearth diameter by this method. Oxygen consumption is estimated at possibly 280 m2 (0.4 t)/t iron made and carbon consumption at possibly 280 kg/t iron. Power consumption for oxygen including auxiliaries is estimated at 304 kWH/t 100 per cent oxygen as 95 to 98 per cent gas. The air separation cycle due to Schuftan is discussed in some detail, and the components of a 1 600 metric ton/day separation plant are briefly discussed. The paper is a theoretical essay.

By A. D. Pienaar

The aqueous chemistry of niobium and tantalum compounds with hydrofluoric acid is a mature field of investigation. In contrast, very little attention has been paid to reactions of these compounds in non-aqueous environments. Similarly, reactions of elemental fluorine with niobium and tantalum compounds have received little more than cursory attention in literature. This work discusses the isothermal reaction kinetics derived from thermogravimetry, using a less redundant approach than is currently the standard. We also discuss the most likely kinetic models for this reaction.

Jan 1, 2012

By G. W. Healy

The smelting of calcium carbide has always been plagued by eruptions, too often with serious injury y or loss of life. To find their causes, the author calculated an internal energy balance using 1992 data for CaC2 and CaO, and older data for the remaining items. The heat content of the charge, and that of the hot gases formed at the reduction temperature, showed that only a fraction of its heat content can be supplied by these gases. Laboratory data on the electrical resistivity of fine mixtures of coke and non-conductive solids were extrapolated to a commercial carbide charge, and then used to calculate the heat produced by current flowing through the mix from the electrode surface to the hearth, which, under favourable conditions, sufficed to bring the charge to 1900 K and higher, at which temperatures reduction could take place.

Jan 1, 1995

By G. T. van Rooyen

In this overview of materials engineering, the author, drawing on his long experience from his student days until the present (emeritus professorship), briefly recounts the main developments in this field: dislocation theory, fracture problems, and the development of alternative materials such as ceramics and polymers. He then takes a look at the future, in which he foresees a greater demand for materials engineers, and leading roles for universities and technikons in collaboration with industry.

Jan 1, 1996

By A. Brodt, I. Mezhibovsky

The traditional vertical transportation systems deliver both gangue and useful mineral. The extraction of the desired minerals requires the construction of mineral processing and waste management plants, driving production costs, and often creating a negative environmental impact. The proposed technology is an integration of mineral processing within the transportation system. The mined rock is crushed and then loaded into a vertical lifting system filled with heavy media, with an intermediate density between the mineral and the waste rock. Driven by gravity separation, the desired mineral floats to the earth's surface, while the waste sinks. Delivered minerals and sunk waste is extracted from the heavy media and is then regenerated to a dry state. The useful mineral should be suitable for further use, and the waste used for backfilling. Remained residues of the heavy media may be recycled and reused in the vertical transportation system. Trials of the proposed process in a pilot format have confirmed economic and environmental advantages over traditional methods of transportation and mineral processing technologies. Keywords: Gravitational processes of mineral processing, separation in heavy media, vertical transportation to earth's surface by floating up

Jan 1, 2020

A manual on mines rescue, safety and gas detection, by J. Strang and P. Mackenzie-Wood. Australia, Austcue, US$25. (Available through Hampo Systems (Pty) Ltd. Reviewer: W. Holding. The book had its origins in a set of notes used by one of the authors for training purposes at the Southern Mines Rescue Station in New South Wales

Jan 1, 1987

COLLOQUIUM AND GENERAL MEETING Developments in gold recovery processes A General Meeting and Metallurgical Colloquium on the above topic were held on Wednesday, 14th March, 1973, at Kelvin House, Johannesburg. Dr J. P. Hugo (President) was in the Chair. The colloquium was attended by 170 delegates, and was opened by the President at 9.15 a.m.

By B. McFadzean, M. Becker, S Geldenhuys, A. Molifie

Mineral hydration is known to result in low flotation recoveries and grades within primary platinum group element ores worldwide. This is because the phyllosilicate minerals derived from hydration reactions of silicate minerals (i) form electrostatic coatings on valuable minerals that in turn hinder collector adsorption, (ii) alter the viscosity of the slurry leading to poor gas dispersion and (iii) decrease the concentrate grade due to naturally-floating gangue. Sodium silicate was investigated because its dispersant, rheology modifying, and depressant properties could prove promising in combating the problems associated with these ore types. Quantitative evaluation by scanning electron microscopy analysis revealed large quantities of serpentine and talc present within the ore sample used in this study, which led to a poor flotation response as indicated by batch flotation tests. Using sodium silicate improved the recoveries and grades at high dosages (>1000 g/t). The zeta potential experiments indicate that improved recoveries are, in part, as a result of the reversal of serpentine surface charge, creating electrostatic repulsion between serpentine and valuable minerals which prevents the coating of valuable minerals by serpentine slimes. Ore dilution and rheology tests indicated that decreased viscosity at high dosages also improved recoveries. Talc micro-flotation tests revealed that the depression of talc occurred at higher sodium silicate dosages, which improved concentrate grade. The processing of nearsurface altered ores is becoming an increasing problem worldwide and the use of sodium silicate proved valuable in mitigating the problems associated with the altered ore investigated in this study. Keywords: Alteration, phyllosilicate slimes, colloidal chemistry

Jan 1, 2020

By M. I. Mabala, B. P. Watson, M. J. Kanda

A failed and stable Upper Group 2 chromitite pillar database was collected in 2007 as part of the PlatMine 1.2 pillar strength project. The database consisted of 177 pillars, of which 33 had failed. Most of the data were collected from various intermediate-depth mines near the town of Thabazimbi. Spans between rows of chain pillars were about 30 m. Pillar loads were estimated using pseudo-3D, elastic, boundary element modelling software. The pillar-load data were evaluated using a maximum likelihood regression analysis to determine strength parameters, following a similar approach to historically acceptable practice. The original formula was named the PlatMine Formula. This paper describes re-evaluations of the database, using a linear formula and several variations of the power formula. The formula providing the lowest standard deviation to the data in the database was considered the best representation of the pillars in the database. Through this rigorous process, a new formula for peak pillar strength was developed for Upper Group 2 chromitite pillars, which is unaffected by volume. The small standard deviation provided by this formula suggests a reliable relationship between strength and w/h ratio, in the w/h ratio range between 1.5 and 4.7. Accuracy, precision, and recall analysis confirmed the efficacy of this equation to the data in the database. Importantly, the new formula provided in the paper does not predict a significantly different strength to the current PlatMine formula, within the confines of the database.

Nov 27, 2024

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 E. M. W. Skinner

Various models of near-surface kimberlite pipe emplacement have been proposed over the years. These include a top-down, phreatomagmatic model, a bottom-up, embryonic pipe model and three top-down explosive dyke models. AU of these models consider kimberlites as essentially the same rock type. However, different kimberlites have different pipe shapes and contain different rock types with very specific mineral assemblages and textures and therefore are likely to have been emplaced by different processes. Some authors have considered local geological differences as the principal reason for the contrasting geology but others argue that, while geological differences might contribute locally the petrographic peculiarities of particular kimberlites may be due mainly to inherent compositional differences specifically in the ratios of juvenile CO2 and H2O.

Jan 1, 2010

By N. Wagner, P. den Hoed, J. D. Steenkamp, S. Soqinase

In pyrometallurgical processes, metal oxides are reduced from molten slag through carbothermic reduction. It is of interest to evaluate the reactivity of the carbonaceous materials towards substances such as slag. Characterization techniques such as coal petrography can provide insight into the influence of feed coal properties and how they potentially dictate reductant performance. This study aimed to compare the petrographically determined organic composition of coal to reductant reactivity. Two South African medium-rank C bituminous coals and one anthracite sample were investigated together with high-carbon ferromanganese industrial slag. The reductant reactivity tests were conducted at 1500°C in a muffle furnace to assess the potential of carbonaceous reductant in reacting with the main slag components. SEM-EDS was applied to understand the extent of MnO (and to a lesser extent, SiO2) reduction from the slag. Coal 2, consisting of a greater proportion of vitrinite (59.5 vol% on a mineral matter-free basis and 54.7 vol% including mineral matter) was the most reactive reductant. The anthracite sample, with the highest inert maceral proportions (71.8 vol% including mineral matter and 76.8 vol% on a mineral matter-free basis), was the least reactive reductant.

Feb 6, 2023

By L. R. Nelson

The efficiency and cost-effectiveness of the aluminothermic reduction process in the production of ferro-alloys and 'pure' metals depends strongly on appropriate balancing of the process energy required. Insufficient energy results in a poor metal yield, while over-supply can result in excessive fuming and possibly even ejection of the reactor contents in a potentially dangerous 'thermite bomb' reaction. This situation becomes even more critical when the suitability of limited quantities of oxide from new sources must be evaluated, dictating smaller-scale testwork. The resulting higher relative heat losses from the system complicates control of the energy balance, and usually requires the addition of thermal boosters or reactant preheating in order to sustain an autothermic process. This often detracts from the quality of the metal product, and adds to the hazards of batch aluminothermic- reduction testing. The development of a small-scale method (producing less than 2,5 kg of metal) for the plasma-assisted aluminothermic reduction of chromium oxide that overcomes such difficulties is described. The technique involves the use of a sealed 50 kVA plasma-arc furnace to provide controlled electrical ignition and supplementary energy to the reactants contained in a water-cooled copper crucible. The application of the method to the evaluation of the aluminothermic production of pure chromium metal from a unique chromic oxide source is reported.

Jan 1, 1996