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By A. J. C. Paterson

High density slurry and paste tailings systems offer significant environmental benefits primarily related to reduced water consumption. However, these potential advantages need to be evaluated on a project-by-project basis. The successful design of a thickened or paste tailings system requires that the system is designed to accommodate the effect of the high density slurry on pumps, pipelines and the overall system operation. The transport system must be designed as an integrated unit and the interaction of all the components need to be considered. The transport system includes the slurry preparation plant, the pump station and the pipeline. This paper assesses the impact of increasing the slurry density in a platinum tailings system from conventional densities to thickened and paste tailings densities. As slurry density and rheology increase, there is a sharp increase in energy consumption for the additional water savings and this additional cost needs to be considered. A case study shows that there is a considerable benefit in increasing densities to the point at which tailings can be still be deposited in the standard manner. For existing systems, increasing density can result in an increased capacity at a marginal power cost, provided that the slurry is not too viscous. For a new system, a smaller pump and pipeline can result in lower capital costs and a marginally higher energy cost, if the optimum density is selected. Keywords: tailings, paste, pumping, slurry

Jan 1, 2004

By F. K. Mulenga, M. Lusambo

"Geometallurgical models are a strategic tool used primarily in planning, diagnosis, and process optimisation. This paper discusses the development of a predictive model for flotation recovery at Kansanshi concentrator in Zambia as a function of key operational parameters. Daily data for milling and flotation from the circuit processing the copper sulphide ore was analysed. This data, recorded over a period of one year of production, was collected from diverse sensors measuring semi-autogenous milling and flotation parameters. Statistical analysis showed that copper head grade, acid-insoluble, and acid-soluble copper, SAG mill throughput, ball mill load, and tonnage of feed material to the SAG mill in the -125 +63 mm size fraction contributed to the prediction power of the proposed empirical model by as much as 90%. Finally, a proposal was made for the development of a simulation model under MODSIM®, a flexible tool for mineral processing applications. This was to test the limitations of the empirical model for future refinement. IntroductionKansanshi is the largest copper mine in Africa and the world’s eighth largest copper producer, with a total of 253 272 t of copper produced in 2016 (Newall, 2017). It is located approximately 10 km north of Solwezi, the capital and administrative headquarters of the North Western Province, Republic of Zambia. The ore deposit consists predominantly of chalcopyrite and malachite with variable amounts of chalcocite, bornite, chrysocolla, and azurite. Because of the varying proportions of sulphide and oxide copper minerals, the ore is classified into three distinct categories: sulphide, mixed, and oxide ores. As such, three separate processing circuits are used to mill the ore down to a standard targeted grind size of 80% passing 150 µm before froth flotation and leaching.A lot of research and development has been done to improve the performance of the sulphide ore circuit at Kansanshi. Despite the research efforts and subsequent operation, the sulphide ore circuit did not attain the target recovery and grade for 2016, primarily because of the complexity of the orebody.The constant drive to achieve the set production target of recovery and concentration grade after froth flotation motivated the development of a tool for the prediction of recovery. The anticipated benefit is not only to enable the plant to apply corrective actions when needed, but also to make a tool available for the quick assessment of the plant performance of the. For that purpose, data spanning June 2016 to May 2017 was analysed, cleaned up and curvefitted to an empirical model. In doing so, key parameters were determined and calibrated to represent the behaviour of the plant."

Nov 1, 2018

Notices Competition for Student Members of the South African Institute of Mining and Metallurgy Each year the Institute offers a prize (or prizes should the entries warrant it) of up to Rl00 for the best paper or dissertation on a topic appropriate to the interests of the Institute. The competition is open to all Student Members of the Institute. NEW PUPLICATION Bougainville-the establishment of a copper mine. St. Kilda (Australia), 'Foundations', 1973. The editors of Foundations, Australia's journal of architecture, engineering, and construction, are preparing the above book for publication late in 1973. The story of Bougainville is told and illustrated NIM REPORTS Report No. 1503 Fast-neutron-activation analysis for the determination of oxygen m rock sam ples. Report No. 1514 The determination by X-ray-fluorescence spectrometry of minor amounts of niobium and tantalum in geological materials. Report No. 1530 AUREOL - an X-ray-absorption analyser for the determination of the uranium on loaded resin in ionexchange plants. Report No. 1534 The determination, by atomicabsorption spectrophotometry, of noble metals in the presence of sodium, barium, or sulphate ions. A comparison of uranium and lanthanum as 'releasing agents.

Jan 10, 1973

By A. K. Verma

Several longwall faces in India have collapsed due to adverse geological conditions and/or inadequate capacity of powered supports. Thus it is imperative to access the capacity of a powered support and its interaction at longwall faces of varying geomining conditions in advance with lots of confidence. To achieve this, an index has been developed to ascertain the front leg pressure (FLP) of chock-shield support and roof-to-floor convergence (RFC) at longwall faces. The index called ?Longwall Face Stability Index? (LFSI) has been developed using statistics and Finite Element Modelling (FEM). The front leg pressure (FLP) and roof-to-floor convergence (RFC) predicted from the LFSI are validated with the field monitored data from two longwall panels. In this study, an attempt is also made to incorporate various geomining conditions into the finite element models typical of longwall mining practice. The index has taken into account the wide variations of geomining parameters like thickness, young?s modulus and friction angle of main roof (small, large), coal type (soft, hard), powered support capacity, and depth of coal seam. Thus, LFSI can be thought of as an index representing combined effect of roof type, geomining conditions and powered support capacity.

Jan 1, 2008

By P. de A. Barbosa, M. G. Bergerman, E. da Fonseca

Located in the state of Pará, Brazil, the Sossego copper and gold mine consists of a heterogeneous geological deposit. Its complexity derives from the hydrothermal alteration, the ore-forming event, which modified the host rocks, such as granite, and metamorphosed volcanic rocks into several mineral assemblages. Different mineral assemblages have led to differences in abrasiveness. A few Bond abrasion tests conducted during the mine feasibility study did not provide a robust abrasiveness database, thus the variation of physical properties for some mineralisation host rocks was not possible. Although several test methods may be used to determine ore abrasiveness, the Bond method is the one most commonly used in the mining industry. Despite being simple to carry out, this test method has certain limitations as to particle size and the amount of ore sample required, which often makes it difficult or even impossible to apply in the early stages of a project and in geometallurgical studies. The Laboratoire Central des Ponts et Chaussées (LCPC) test may provide an alternative, as it requires a smaller amount of fine-grained material. This study aimed to provide data on wear properties of various ore types from Sossego mine using the LCPC method. It included validation for 40 different ore samples using the Bond test as a benchmark. The results showed good correlation, indicating it may be possible to use the LCPC test as an alternative to the Bond abrasion test. A high abrasiveness index was found for acid metavolcanic rock, confirming the results observed in Sossego mill’s operating database, pointing to a mineralogical drive to amorphous quartz. Keywords: Abrasiveness, geometallurgy, copper ore, Bond abrasion test, LCPC

Jan 1, 2020

By C. Liu, G. Zhu, Y. Liao

"The effect of particle size on the flotation kinetics of a low-rank coal was compared using air bubbles and oily bubbles. Five kinetic models were applied to the data from the tests to derive the relationships between the flotation rate constant, the ultimate recovery, and the particle size fraction. The results show that oily bubble flotation (OBF) resulted in a lower ash content and higher combustible recovery than air bubble flotation (ABF), especially for the -125 ??m size fraction. It was found that the flotation of a low-rank coal in individual size fractions using air bubbles and oily bubbles can be best described by a first-order model with rectangular distribution of tendency to float and the classical first-order model respectively. OBF yielded higher ultimate combustible and ultimate ash recoveries. The flotation rate constants for OBF were higher in all the models except for the fully mixed reactor model. The ultimate recovery values and modified rate constants first increased and then decreased as the size fraction decreased. The trend of the flotation rate constants as a function of the size fraction was similar in all the models except for the fully mixed reactor model. IntroductionFroth flotation is a physical-chemical separation method that is widely used in coal beneficiation for the separation of fine coal particles from associated minerals in water slurries. Since the cumulative recovery of a component in the concentrate is proportional to flotation time, the flotation process can be considered as a time-rate recovery process (Emad, Mohammad, and Bahram, 2010). Therefore, the general mathematical flotation models that incorporate both a recovery and a rate function can describe flotation time recovery profiles.The kinetics of coal flotation, as a measure of coal recovery as a function of time, have been studied in detail by many researchers. Different kinetic models have been proposed and published in the literature. The order and the rate constant are strongly dependent on the flotation conditions, such as the particle size and size distribution, the bubble concentration, the reagent concentrations, and the operating parameters (Zhang et al., 2013). The particle size has been reported to affect the flotation kinetics to a great extent (Yoon, Luttrell, and Asmatulu, 2002; Ntengwe and Witika, 2011). Numerous researchers have studied the aspects of flotation kinetics while paying special attention to particle size (Meftuni and Ibrahim, 1997; Rubinstein and Samygin, 1998; Graeme, 2012). However, the effect of particle size on the flotation kinetics of a low-rank coal has been relatively little discussed."

Jan 1, 2017

By H. Zhong, S. Wang, Q. Sun

The flotation performance of 2-(benzylthio)-acetohydroxamic acid (BTHA) to diaspore was first evaluated by micro-flotation tests. The adsorption mechanism was further investigated through adsorption experiments, molecular dynamics simulation, zeta potential and X-ray photoelectron spectroscopy (XPS). The results demonstrated that BTHA had a stronger ability to collect diaspore in comparison with another diaspore collector, benzohydroxamic acid (BHA), especially under the condition of weak alkalinity, the flotation recovery can reach 90%. The more negative interaction energy (ΔE) between BTHA and diaspore from molecular dynamics simulation results revealed that BTHA had an easier adsorption ablity onto diaspore than BHA, which matched with adsorption results. The zeta potential and XPS revealed that Al atoms might bond with -C(=O)NHOH by formation of Al-O bonds. Keywords: Diaspore, flotation; adsorption; 2-(benzylthio)-acetohydroxamic acid

Jan 1, 2020

By J. H. Bulloch

A study of fatigue-crack growth in a rail steel and an associated weld metal showed that fatigue cracks in the deformed rail steel had slightly higher growth rates than in undeformed rail steel. The crack growth in the weld metal was appreciably faster than either and exhibited plateaux that were above the upper bound reported for rail steels, The cyclic-life curves showed that, over an indicated service period of 8 years, both the deformed and the undeformed rail steels should not be prone to fatigue failures under the defined mechanical conditions, For the weld metal, the cyclic-life curves indicated the possibility of premature fatigue failures but, as al the input parameters represented a 'worst case' situation, this assessment could wel be ultra-conservative. In the rail steels at low levels, the extension of the fatigue cracks occurred by ductile striated growth. At a Kmax level of 35 MPa .J/n, cleavage facets formed initially on the surface of the fatigue fractures, and their incidence increased with Kmax until the onset of fast failure, The microstructure of the deformed rail steel exhibited cleavage facets at a slightly lower Kmaxvalue, and contained more cleavage facets at a given K level, than the microstructure of the undeformed rail steel. The average size of the cleavage facets agreed well with the pearlite nodule dimensions of 60 to 100 ILm. The microstructure of the weld metal was much coarser and contained columnar grains. At al f.K levels, the dominant fracture mode was cleavage, with small isolated regions of ductile striated fatigue cracks. The size of the cleavage facets varied from 150 to 600 ILm, and such a variation may be explained by the fact that cracking generally occurs in association with the pro-eutectoid ferite phase, and that this phase is distributed along some boundaries of prior austenite grains and the boundaries of the coarser columnar grains.

Jan 1, 1987

By M. de Paiva Bueno, G. Barakos, S. Luukkanen, K. Luolavirta

Lifelong development of skills and knowledge of mining professionals has been recognised worldwide as a challenge for the development of a sustainable mining sector. Practical training and learning over the mine value chain are among the key components of this knowledge building. However, the number of available mining sites for continuous education in real-life conditions is limited; also the specifications for practical training are not clearly defined. The project MINETRAIN was developed to (i) evaluate the potential of the shortly-closing Pyhäsalmi Mine, in Finland, as an educational and training underground facility, and thus, prolong the utility of the existing modern mine site, and to (ii) develop and test practical training programs for mining and mining industry related professionals. Pyhäsalmi facilities enable training among all disciplines related to the overall mine value chain and hence, the project has potential in providing education for sustainable management of any phase of mining operations. The philosophy underpinning the MINETRAIN approach to mining activities is to assist the clients in minimising the technical and financial risks by providing a competitive and skilled personnel asset while optimising the value drivers in their operations. As a first major step of the project, a pilot training course, involving basic-level online theory and hands-on training at the Pyhäsalmi mine site, was undertaken in summer 2019. Feedback from the pilot test group of mining professionals has proved to be very positive. The comments from the participants strongly emphasised the added value of the practical onsite training for improved understanding in the large scale picture of mining on the entire value chain. As a second phase of the project, a new test course was launched in June 2020. The second MINETRAIN course focuses on new technologies and digitisation in mining. Keywords: Lifelong learning, mining education, mine value chain, research and education mine, educational research project

Jan 1, 2020

By A. V. Wilke

The metallurgical balance is a tool usually applied in the mining industry. The final result is used for different objectives. All the existing software for the calculation of the metallurgical balance fits the numbers in such way that a logical correspondence between the input and output materials exist. Nevertheless, the accuracy of the metallurgical balance strongly depends on the quality of the input data. Division El Teniente of Codelco Chile has been optimizing the metallurgical balance system the last 14 years. As well, it has participated in the development of corporate norms to optimize the operations of sampling and weighting systems. As a consequence, the precision and the accuracy of the metallurgical balance have been improved in a significant way. Also the understanding of process variability was enhanced. The improvements have allowed Codelco to discover and minimize hidden losses and to better support top management decisions. The present paper describes the main aspect of the continuous improvement process.

Jan 1, 2009

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

Mr R. C. J. Goode (President) was in the Chair. There were also present one hundred and thirty-six Members, including Messrs R. J. Adamson, Dr J. M. Bereza, Messrs M. Barcza, H. Britten, H. E. Cross, J. K. E. Douglas, G. W. Holl, Dr J. P. Hugo, Messrs C. J. Irving, P. Lambooy, Prof J. de V. Lambrechts, Mr D. G. Maxwell, Dr J. T. McIntyre, Prof R. P. Plewman, Messrs J. F. Reid, V. C. Robinson, H. Simon and P. W. J. van Rensburg (Council Members), G. Abdinor, C. B. Anderson, H. F. Allan, E. L. Atkinson, V. C. Barnes, J. F. Bennet, W. W. Bath, R. E. Barnes, C. L. Berrange, J. P. Buds, Dr W. Bleloch, Dr L. A. Bushell, Messrs D. L. Carson, D. D. Cornier, R. S. Cooke, R. R. M. Cousens, J. P. Coetzee, W. L. Cousens, J. Daniel, R. Daniel, D. G. Davies, S. K. de Kok, V. S. Dillon, B. S. Dykhouse, Dr H. G. Denkhaus, Dr R. S. J. du Toit, Messrs J. I. L. Edwards, R. C. Espley-Jones, J. C. Ferguson, D. T. Fourie, A. Frost, W. R. Flook, Dr N. P. Finkelstein, Messrs D. F. H. Grave, J. M. Gosnell, E. K. Gibson, R. W. Grout, E. W. Gray, J. S. Hay, F. G. Hill, H. N. Hepker, J. P. Hoffman, C. G. Hodges, D. M. Jamieson, Dr D. G. Krige, Dr N. P. Lategan, Messrs J. E. Laschinger, J. Levin, C. P. Linsell, J. O. Kempe, F. Q. P. Leiding, A. I. Lee, D. G. Malan, W. Malan, H. W. McCulloch, R. D. R. Munro, C. S. Macphail, Dr T. F. Muller, Dr A. J. Nel, Messrs S. C. Newman, G. G. Nicolson, G. Y. Nisbet, J. A. Papenfus, E. Pavitt, W. H. Pittie, P. C. Pirow, P. G. D. Pretorius, N. F. Peverett, R. P. Plaskett, M. P. Pearse, W. Pilkington, T. K. Prentice, Dr R. E. Robinson, Messrs R. B. Richardson, R. St. J. Rowland, R. J. Rouillard, F. O. Read, W. T. Ruhmer, D. Savile Davis, E. F. Statham, K. Smith, H. Simon, G. B. Stern, C. D. Storrar, Dr N. B. Strydom, Messrs J. H. Talbot, A. H. Taute, J. H. Taylor, G. W. Thiel, K. A. Tweedie, W. E. Vance, Dr H. J. van Eck, Messrs L. W. P. v. d. Bosch, P. A. von Wielligh, N. B. Vinson, S. J. Venning, W. M. Walker, D. T. Watt, Dr H. J. R. Way, W. T. L. Wayman, R. G. Williams, L. G. S. Wright and C. F. Ziervogel. Thirty Associate Members, including Messrs E. Collier, G. Douglas, D. H. Gray, H. G. Gericke, R. Hemp, N. M. Hayne, A. Hazell, P. M. Johnston, D. E. Love, Miss V. E. Marting, Messrs R. G. Matthews, P. Malpage, D. Nicolle, R. Pendreigh, J. J. Rosslee, R. W. Shearer, R. F. Taylor, G. C. Thompson and D. A. Viljoen. Three graduates, Messrs G. S. Giles, J. P. Hatfield and J. A. Jones. Five Affiliates, Messrs A. C. Collie, E. Collier, A. B. Cook, W. Jacobs and G. M. Thain. Two Students, Messrs C. J. Bircher and R. P. Henrard.

By K McLeod

{Figure 1} The Placer Dome-Western Areas Joint Venture?s South Deep Twin Shaft Complex, has been planned and dreamt of for nearly 20 years. [ ] The gold reserve at the bottom of these shafts is one of the largest single intact gold deposits remaining in the world today, containing an estimated 57 million ounces. This is about the size of the fictitious reserve created by inventive prospectors for the infamous BREEX scam in Kalimantan about ten years ago and about the size of the reserve Harmony claims as its own in it?s current hostile takeover bid for Gold Fields. So it is certainly of the magnitude that dreams are made of. [ ] {Figure 2} The unique orebody lies in a continuous, highly concentrated, thick deposit, an estimated twelve kilometres along dip and a two kilometre wide strike. Due to the thickness of some sections of the reef, up to 60 metres in places, bulk mining methods will be required for it?s extraction. [ ] {Figure 3} The twin shaft system comprises a main shaft which had an initial design depth of 2765m and a ventilation shaft of 2760m, making them the world?s deepest to be sunk in a single lift. The shafts were

Jan 1, 2005

By N. Snyders, I. Mwandawande, S. M. Bradshaw, M. Karimi

"Computational fluid dynamics (CFD) was applied to predict the average gas holdup and the axial gas holdup variation in a 13.5 m high cylindrical column 0.91 m diameter. The column was operating in batch mode. A Eulerian-Eulerian multiphase approach with appropriate interphase momentum exchange terms was applied to simulate the gas-liquid flow inside the column. Turbulence in the continuous phase was modelled using the k-_ realizable turbulence model. The predicted average gas holdup values were in good agreement with experimental data. The axial gas holdup prediction was generally good for the middle and top parts of the column, but was over-predicted for the bottom part of the column. Bubble velocity profiles were observed in which the axial velocity of the air bubbles decreased with height in the column. This may be related to the upward increase in gas holdup in the column. Simulations were also conducted to compare the gas holdup predicted with the universal, the Schiller-Naumann, and the Morsi-Alexander drag models. The gas holdup predictions for the three drag models were not significantly different. IntroductionColumn flotation is an important concentration technology that is used in the mineral processing and coal beneficiation industries. The growing interest in the use of column flotation in mineral processing has been attributed to the simpler flotation circuits and improved metallurgical performance compared to conventional flotation cells (Finch and Dobby, 1990). Flotation columns have also found other applications outside mineral processing, such as de-inking of recycled paper (Finch and Hardie, 1999).In column flotation, a rising swarm of air bubbles generated by means of air spargers is employed to collect the valuable mineral particles and separate them from the gangue minerals in a countercurrent process. Wash water, which is continuously fed at the top of the column, is used to eliminate entrained particles and stabilize the froth. The column volume can be divided into two sections – the collection zone in which the bubbles collect the floatable mineral particles, and the cleaning zone (or froth zone) where product upgrading is enhanced through the removal of unwanted particles entrained in the water mixed with bubbles from the collection zone."

Jan 1, 2019

By K. Wu, X. Yuan, R. Du, C. Chao

"Lump coals are currently employed to replace part of coke in the COREX melter-gasifier for ironmaking. In order to optimize the process, the pyrolysis kinetics of Datong coal used in the COREX process was studied. A sectioning method is proposed to overcome certain shortcomings of traditional methods for studying the pyrolysis of coal. Based on the sectioning method and the characteristics of the coal pyrolysis process, the temperature range was divided into three intervals. In the first, the adsorbed gas and water of crystallization were removed; in the second, the coal was transformed into semi-coke; and finally, the semi-coke was further coked. Three models of reaction were then established to study the different stages of the pyrolysis process: an interfacial chemical reaction model, a random pore model, and internal diffusion with shrinking volume. The results showed that the experimental data fitted the models well, with all correlation coefficients (r2) exceeding 0.97. Finally, the mechanisms of reaction and kinetic parameters for each temperature range were derived, which laid a necessary foundation for optimizing the usage of non-coking coal in the COREX process. IntroductionIn view of environmental concerns and the shortage of good-quality coking coal, the COREX process was designed as a new ironmaking technology, attracting wide attention (Fang et al., 2005). In this process, lump coal and a small quantity of coke are directly charged into the melter-gasifier, where a high-quality reducing gas is generated for the reduction of the iron ore (Kumar et al., 2009). In contrast to the conventional blast furnace (BF) process, non-coking lump coal is used as the reducing agent and as the energy source, and the iron charge consists of lump ore and/or pellets, which results in the elimination of the coke oven and sintering plant as well as a decrease in investment (Kumar et al., 2009; Liu et al., 2012).Although the COREX process has made great progress, in current production practice the energy consumption greatly exceeds that of the BF, and it requires a certain amount of coke to maintain permeability of the semi-coke bed (Wang et al., 2008). In addition, the process consumes a lot of good-quality lump ore and lump coal. The pyrolysis of coal is the first step in most coal conversion processes, such as carbonization, gasification, and combustion (Aboyade et al., 2013). The reaction behaviour of lump coals in the COREX melter-gasifier is crucial, as it determines energy utilization and gas composition, as well as the metallurgical properties and particle size of the semi-coke produced (Prachethan et al., 2011; Zhang et al., 2014). In order to optimize the process and improve its competitiveness, the pyrolysis kinetics of Datong coal used in Baosteel’s COREX C-3000 facility was studied."

Jan 1, 2016

By D. K. Chibwe, P. Taskinen

"Typical operation of industrial Peirce-Smith converters (PSCs) used in copper smelting results in several common phenomena such as excessive splashing, slopping, tuyere blockage, and low converter campaign life due to high tuyere line refractory erosion. These phenomena give rise to an inefficient operation from both a process and an energy perspective.This investigation presents numerical and physical models of flow patterns, mixing, solid-liquid mass transfer, and slag-matte phase distribution in a slice model of an industrial PSC. For physical simulations, a 0.2-scale cold model was developed. Water, kerosene, air, and sintered benzoic acid compacts were used to simulate matte, slag, injected gas, and solid additions into the PSC. The two- and three-dimensional numerical simulations of the three-phase system were carried out using volume of fluid (VOF) and realizable k- e (RKE) turbulence models to account for the multiphase and turbulence nature of the flow respectively. These models were implemented using the commercial computational fluid dynamics numerical code FLUENT.Both numerical and physical simulations were able to predict, in agreement, the mixing and dispersion characteristics of the system in relation to various blowing conditions employed in the investigation. Measurement of mass transfer characteristics conclusively indicated that fluid flow in the PSC is stratified. Both blowing configuration and volume of slag in the converters were found to have significant effects on mixing propagation, wave formation, and splashing.The splashing and wave motion in these converters are known to cause losses of metal/matte and potential production time due to the requirement for intermittent cleaning of the converter mouth, resulting in reduced process throughput. To prevent these losses and hence to increase process efficiency, we propose a combined blowing configuration using top lance and lateral nozzles. The numerical simulations were conducted on combined, as well as lateral, blowing conditions. In comparison, the recent results of the combined blowing concept were found to be encouraging."

Jan 1, 2014

By P. Vlasak, Z. Chara

Since the flow behaviour of slurry containing both coarse and very fine particles has not been sufficiently clarified, the paper focuses on the experimental study of the particle size distribution, time and intensity of shearing and the effect of chemical agents on the hydraulic gradient versus the average velocity relationship of fluidic ash slurries in the laminar and turbulent regimes. The recirculation pipe loop with smooth stainless steel pipes of inner diameters 26.8 and 36 mm was used for measuring the ash-water mixtures flow parameters. The dense fluidic ash-water mixtures behave as time dependent non-Newtonian slurries. Shearing, the addition of bottom ash or chemical agents, favourably affects the flow behaviour of dense fluidic ash-water slurries. The hydraulic gradient in the laminar region significantly decreases; also the laminar/turbulent transition region is reached for lower flow velocity.

Jan 1, 2007

By C. A. Snyders

The research presented in this paper investigated the practical aspects of the recovery of platinum group metals (PGMs) from a dilute cyanide leach solution containing base metals, in a manner similar to that used for gold extraction in a typical CIP process, and focuses on both the adsorption and elution stages. The carrier-phase extraction of precious metals using activated carbon offers significant advantages over other processes in terms of simplicity, the high pre-concentration factor, rapid phase separation, and relatively low capital and operating costs. As a sorbent, activated carbon is still by far the most important material because of its large surface area, high adsorption capacity, porous structure, negligible environmental toxicity, low cost, and high purity standards. Adsorption tests were conducted on a pregnant alkaline leach solution (0.15 ppm Pt, 0.38 ppm Pd, 0.1 ppm Au) resulting from cyanide extraction performed in column leach tests. The initial adsorption rates of Pt, Pd, and Au were very fast and recoveries of these three metals were approximately 90 per cent after 2 hours, and 100 per cent for Pt, 97.4 per cent for Pd, and 99.9 per cent for Au after 72 hours. The parameters that influence the extraction of PGMs and Au were examined to assess their relative importance during the adsorption process in order to provide the basis for process optimization. The concentration of thiocyanate was not identified as significant factor for PGMs adsorption, while Ni concentration was the most significant extraction process parameter. Base metal cyanide complexes adsorb and compete with the PGM complexes for sites on activated carbon, and while copper adsorption can be minimized by adjusting the residence time, Ni adsorbs at approximately the same rate as the PGMs, influencing the loading capacity and adsorption kinetics of the PGMs. The feasibility of eluting platinum and palladium cyanide complexes from activated carbon using an analogue of the AARL process was investigated. Platinum and palladium elute from activated carbon almost to completion in 4 to 5 bed volumes at 80°C, while the elution of gold at this temperature is slow, with a significant amount of gold still to be eluted after 16 bed volumes. The equilibrium loading of gold is exothermic in nature (Fleming and Nicol, 1984) which will result in an increase in gold elution kinetics with an increase in temperature at similar pre-treatment conditions. A similar result was found for the elution of Pt and Pd. Cyanide pre-treatment was found to have a significant influence on PGM elution. Higher cyanide concentration in the pre-treatment step results in more efficient elution up to a point, and results suggest the possibility of an optimum cyanide concentration, beyond which elution efficiency starts decreasing due to increased ionic strength.

Jan 1, 2013

By A. Melkumyan, K. Silversides

Chemical assays collected from blast holes have the potential to greatly improve the local accuracy of grade models. Assays must be assigned to the correct domain or they will introduce errors. Boundary surfaces created from the widely spaced exploration holes (~50m) do not provide sufficient resolution to correctly delineate domain boundaries between closely spaced (3-5m) blast holes. This paper demonstrates a probabilistic method for classifying production blast holes near domain boundaries. Gaussian processes models are applied to identify the assays and assay ratios most suited to distinguishing particular domains. These models are combined with the exploration based surfaces to classify the blast holes that can be confidently assigned to a particular domain. A reiterative method is used, where information from the blast holes closest to the exploration holes is used to label blast holes further away. This is tested on two domain boundaries in a banded iron formation-hosted iron ore mine.

Sep 1, 2021

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