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By C. O’Donova, E. Adam, L. A. Torres-Cruz

High-profile failures of tailings dams illustrate the significant safety challenge they pose. The decant pond of such facilities plays an important role in their mechanical stability. This study investigates the feasibility of using public multispectral data collected by the Sentinel-2 satellite mission to monitor the area of the decant pond. The Sentinel-2 mission has a revisit time of 5 days and the data used herein has a spatial resolution of 10 m/pixel. We considered two water-detection metrics: the normalized difference water index (NDWI) and the near-infrared (NIR) reflectance. Implementation of the method at a South African platinum tailings dam illustrates the capabilities and limitations. Public, sub-metre resolution true-colour imagery served as validation data. There is good agreement between pond area measurements based on Sentinel-2 and the validation imagery. These results suggest the feasibility of using Sentinel-2 data to supplement the monitoring of the decant pond.

Apr 2, 2022

By Ri McCrindle

"Analytical Chemistry is moving rapidly towards a concept where technicians have to do no more than press a button and print the result from an instrument that is a featureless ""black box"". In this environment we must ask how much traditional ''wet chemistry"" is still required for undergraduate students.Methods that can be fully automated will become increasingly important. An example of such a method is that of the slurry nebulisation and ICP-OES analysis of precious metal concentrate. Grinding the sample and adding Triton X100 as dispersant, complete analysis time is in the order of 90 min as apposed to the traditional Carius tube method which could take as long as 24 h. For most of the elements studied, the results using ""in-house"" reference material for calibration and as samples as well were within the 95% confidence level. Elements of low concentration in the sample (As, Bi and Se) were not within that level."

Jan 1, 2014

The Annual General Meeting of the South African Institute of Mining and Metallurgy was held in Kelvin House on Wednesday, 23rd August, 1978. Mr P. A. von Wielligh (President) was in the chair and declared the meeting open at 16h00. Obituaries The President: It is my sad duty to announce the death of the following members of the Institute : T. K. Prentice and W. E. Gooday, Honorary Life Fellows and Past Presidents; B. S. Allen, J. A. Becking, L. H. Mellor, C. A. Pratt and W. T. Stobart, Life Fellows; J. P. Burls, E. Cameron, M. L. Fitzgerald, C. G. Hinds, A. J. Patrick, H. O. Reisener, N. D. Shuttleworth, N. Soldan and J. Whitehouse, Fellows; T. Gemmell, J. S. A. Oelofse, W. D. Rowe, R. L. J. Shearman and G. M. Thain, Associates. As a mark of respect to the memory of the deceased and in sympathy with the bereaved, I ask you to rise and observe a few moments' silence. Welcome The President: It is now my very pleasant duty to welcome our distinguished guests and visitors. We are pleased to have with us Mr D. A. Etheredge, Vice¬President of the Chamber of Mines of South Africa, and Mrs Etheredge. Mr Etheredge this afternoon represents the mining industry at our meeting. The importance of the close link that over the years has been maintained between the industry and our Institute speaks for itself. We are, therefore, grateful that he and Mrs Etheredge have managed to find the time to be with us this after¬noon.

Jan 1, 1979

By A. Mehmet

The feasibility of employing a multi-stage column for the recovery of gold from cyanided pulps using extruded carbon was investigated on a pilot-plant scale. It was anticipated that the column contactor would have certain advantages over conventional carbon-in-pulp plants. The feed pulp was screened to remove particles larger than 200 µm and was diluted to a solids concentration of about 23 per cent. The column was operated continuously for 24 days at a pulp flowrate of about 60 l/min and a carbon flowrate of 44 litres per day. The average gold concentrations in the barren pulp 12 hours and 24 hours after the carbon transfer were 0,012 and 0,031 mg/l respectively. The average gold concentration in the solution phase in the feed was about 2,0 mg/L. The average gold Loading on the carbon was 6290 g/t. Calcium carbonate was removed from the loaded carbon by treatment with 0,5 bed volumes of 2,8 per cent hydrochloric acid solution at ambient temperature. The calcium concentration was reduced from about 17 000 to about 500 g/t The carbon was eluted by the Zadra procedure at atmospheric pressure. The results of the elution were highly satisfactory, the gold concentration being reduced from about 6000 to about 20 g/t in 48 hours. The eluted carbon was regenerated in a Rintoul kiln at about 700°C. The activity of the regenerated carbon was found to be at least as high as that of fresh carbon.

Jan 1, 1992

By D. G. Davies

Colloquium on the construction of Slimes dams The following contributions to the discussion at the above Colloquium were received for publication. D. G. DAVIES* This contribution is really a follow up on Mr H. E. Cross's paper on the conversion of a uranium plant to the recovery of zinc (published in the Journal, November 1973). Ecological aspects of slimes-dam construction by H. T. CLAUSEN C. M. VAN STADEN* Published in the Journal, December 1973 I should like to thank Mr Clausen for his paper and to congratulate the Research Laboratories and the Vegetation Unit of the Chamber of Mines on the excellent work they have done in the past, are still doing at present, and, we hope, will continue to do in future, on the testing of various methods for the covering of slimes dams and specifically on the vegetation of dams. The impact of slimes-dam formation on water quality and pollution by R. T. RUDD* C. M. VAN STADEN Published in the Journal, December 1973 I should like to congratulate Mr Rudd on the paper he presented. Congratulations are also due to the Department of Water Affairs for the practical attitude they have taken in implementing the Water Act. Slimes-dam construction in the gold mines of the Anglo American Group by W. T. RUHMER W. R. FLOOK* Published in the Journal, February 1973 Methods for the construction and operation of slimes dams in the gold-mining industry do not appear to have changed much for some considerable number of years.

Jan 3, 1974

Publications of the 13th CMMI Congress, by J.T. Woodcock (editor-in-chief). Melbourne, Australasian Institute of Mining and Metallurgy, 1986. Set $A90,00. Reviewer: J.D. Austin

Jan 1, 1988

By M. S. Shishvan, S. Klerkx, J. Sattarvand

The inaccurately drilled blastholes from the planned locations cause changes in the spatial distribution of explosive energy; some areas of the blast may receive a surplus of explosive energy, while others see a reduction. This results in less uniform fragmentation, reduces the efficiency of mining operations and increases costs. Additionally, it can lead to slope stability issues, flyrock, and uneven benches. To mitigate these problems, this study attempts to reduce differences in the explosive energy distribution (EED) through an optimization approach that adjusts the height of explosive columns in blastholes. Two optimization algorithms, the Genetic Algorithm (GA) and Tabu Search (TS), are tested. The latter is preferred for its local search neighbourhood strategy that gives superior computation times. Applications on different blast patterns result in improvements of 0.53–1.54%, or 2.14–3.94% when only the affected blocks are considered.

Sep 1, 2021

By T. Könighofer

Processing of copper-cobalt orebodies in Africa?s Copper Belt has received much recent attention with various flowsheet options being developed specifically for the refining of cobalt. The upfront leach and copper refining routes are well understood whereas the refining of cobalt is more complex, requiring numerous impurity removal steps to produce high purity metal. This paper describes an investigation into one of the possible processing steps in the refining of cobalt, namely, cobalt solvent extraction. The paper emphasizes the importance of upstream impurity removal to achieve the required cobalt solvent extraction feed composition. 18% (v/v) Cyanex 272 in an aliphatic diluent is used as the organic phase. The pH profiles in the various stages are evaluated in order to obtain a raffinate containing <10 mg/[l] cobalt and with a low magnesium content reporting to the stripping section. The pH profile is also used to minimize the impurity deportment to the cobalt electrolyte. The effect of adding a preneutralization stage before feeding the organic phase to the extraction circuit is investigated. Zinc build-up in the stripping stages is also looked at. It is recommended for the purpose of this study that four extraction stages at an O:A ratio of 1 with a pH profile from 4.9 to 5.7 be employed. The pH is controlled with 40 g/[l] NaOH. A preneutralization stage is required, where the organic phase is contacted with 10 M NaOH. Two scrubbing stages are recommended at an O:A ratio of 40, using the cobalt electrowinning advance electrolyte as the scrub liquor. Three strip stages are to be employed at an O:A ratio of 0.67.

Jan 1, 2009

By D. J. Briggs

The use of backfill for regional and local support is receiving increasing attention in the South African mining industry. To ensure that fill materials meet the purposes for which they are intended, a test procedure to characterize the load-deformation characteristics for a particular fill was developed. The major variables identified as influencing the load-bearing proper- ties of backfill include particle size distribution, porosity, width-to-height ratio, strain rate, confinement, water content, and additives. The influence of these variables on the performance of backfill is described in relation to crushed waste, dewatered tailings and deslimed tailings. These investigations confirmed that no single test could be used to define the overall stress-strain behaviour of a selected fill. Nevertheless, resulting from this work, a preliminary standard test procedure for uncemented backfill materials, com- prising a suite of tests in unconfined and confined compression, has been formulated. This test procedure is a significant advance towards a more meaningful evaluation of fill materials. Test results will also provide numerical information for input use in computer programs for the design of mine layouts incorporating backfill.

Jan 1, 1988

By T. Szendrei, S. Tose

The generally accepted view in rock blasting is that the sources of energy for the fracture and movement of rock reside in the shock wave and gas action resulting from the explosion, and yet the mechanisms by which these sources interact with the rock have remained unclarified. It has also been noted that up to 50% of the work capacity of an explosive released in a blast cannot be accounted for by field measurements of energy partitioning. In this study, we describe a physical model that details the response of rock to both shock wave and gas action. An analytical model based on momentum conservation is derived to describe the dynamics of shock-driven expansion of the blasthole. Radial expansion of the hole is the key parameter that permits the derivation of the following characteristics of rock response to shock loads: hole expansion time; volume of displaced rock; energy consumed per unit volume; expansion energy efficiency; stress wave pulse length; gas pressure in enlarged hole. Soon after the completion of hole expansion, the shock wave degenerates to an elastic stress wave that runs through the burden. Blasthole expansions of between 50% and 300% of diameter are completed in under 1 ms and, depending on rock properties, consume 32% to 42% of the detonation energy or about 55% of the available mechanical (Gurney) energy. Gas pressure in the enlarged holes in five rock types is between 35 MPa and 650 MPa, and drives the mass movement of burden rock.

Oct 9, 2024

By A. S. Geldenhuys, L. S. Leal Filho, D. A. Deglon, T. C. Souza Pinto

The Brucutu iron ore mine (Minas Gerais, Brazil), is Vale`s largest iron producing operation achieving around 21 million tons by flotation of iron ore pellet feed per annum. The run of mine is mainly composed of hematite (70%) and quartz (30%). After preparation (comminution and sizing), the slurry is submitted to reverse cationic flotation of quartz under basic conditions (pH=10.5). Corn starch (500g/t) is used to depress Fe-bearing minerals and alkyl (C10) ether amine is the collector for quartz. The flotation circuit is arranged in a “coarse” section utilizing dated 28m3 square Wemco cells, and a “fine” section utilizing modern 160m3 tank cells. The “coarse” flotation circuit treats material from the deslime cyclone underflow while the “fine” circuit treats material from the deslime cyclone overflow and the pulp product from the “coarse” circuit scavenger 2 bank. Evaluation of process performance is of high importance as even small gains can lead to large improvements for this high throughput plant. Cell-by-cell samples of the froth products, selected feed and pulp-products were analyzed for flow-rate, overall chemical composition, and particle size distribution. Bank-by-bank samples of the feed, froth and pulp products were analyzed on an assay-by-size basis. In addition, the flotation cells were hydrodynamically characterized and gas dispersion parameters assessed. The detailed experimental program resulted in a thorough characterization of the process and enough information to produce a calibrated floatability component model. Longer mainline residence time results in significant Fe2O3 losses while yielding little benefit in terms of SiO2 grade. Scavenger 2 has twice the residence time of scavenger 1 while having to treat only 10% of the SiO2; resulting in high Fe2O3 recoveries to the froth and poor separation. In addition, it is shown that the Fe2O3 exhibits true flotation behavior resulting in increased Fe2O3 losses. The detailed data collection campaign coupled with in depth data analysis and modelling gives unprecedented information and highlights avenues for optimization that was previously unknown. Keywords: Froth Flotation, Process modelling, Process evaluation, Metallurgical survey

Jan 1, 2020

By Y. Belloa, C. Dorfling

The presence of impurities such as Se and Te in copper sulfate leach solutions can have a detrimental effect on electrowinning efficiency. These elements can be removed from the solution by precipitation; sulfurous acid is commonly used as a precipitation reagent, while studies have also reported improved precipitation with the addition of metallic copper. This paper aimed to determine the effect of conditions used during the solution purification step on the characteristics of the precipitates that formed. Precipitation tests were performed in a batch reactor with copper sulfate leach solutions obtained from the base metal refinery of a platinum group metal producer. The effects of agitation rate (250 and 500 rpm), SO2 flow rate (4.4 and 5.8 mL/min.L), temperature (65 and 95°C) and metallic copper powder addition (0, 1 and 2 g/L) on the characteristics of the precipitates formed during the tests were determined using scanning electron microscopy. The major phases detected in the precipitates formed when using SO2 as the only precipitation reagent were elemental Se and Cu2Se. The introduction of copper powder led to the production of different sulfide phases for the experiments conducted at 65°C irrespective of the values of the other reaction parameters. At 95°C, Cu crystals, Cu2O and tiny inclusions of Cu2(Se,Te) were the prominent phases noticed in the precipitates produced from all the tests conducted with copper powder as precipitation enhancing reagent. Keywords: copper, precipitation, selenium, tellurium

Jan 1, 2020

By M. Paricheh, M. Rahmanpour

"The cost efficiency and high reliability of semi-mobile or combined in-pit crushing-conveying (IPCC) and truck systems, compared to conventional truck-shovel systems alone, make them more attractive for use in modern mining operations. Semi-mobile systems, which are the most common systems in open-pit mining, combine the advantages of both systems such that fewer trucks are required, operating costs are lowered, and the operation is environmentally sound. One major aspect of utilizing this system is determining the location of the in-pit crusher and the timing of relocations. Facility location models are rarely used in the mining sector, but these models may be used for solving the in-pit crusher location problem. To this end, the main parameters affecting the IPCC location in open pit mines are reviewed and examples given of how they are applied to a dynamic location-relocation problem. Subsequently, the model is implemented for Sungun copper mine in Iran and the number and the exact time of relocations of the IPCC units determined. IntroductionNowadays, compared to the last century, metal prices are more volatile, average and cut-off grades have decreased, and stripping ratios increased. From a mining engineering point of view, many ore deposits with simple access and low capital and operating costs have been exhausted. The remaining deposits are located at depth, with no easy access, which requires high initial capital costs (Osanloo, 2012). Mining at great depths and high tonnages is possible only through open pit mining. Today, more than 80% of open pit mines in the world use a shovel-truck system for loading-hauling operations (Osanloo, 2010).With increasing depth, haulage distances increase and the number of loads per truck decreases. As a result, fuel, tyre, and depreciation costs per ton increase. Owing to these factors, open pit mining at great depth (300 to 1000 m) using shovel-truck systems faces some technical and economic problems. In-pit crushing-conveying (IPCC) systems have been known in the mining industry for many decades. The idea was introduced in 1956 in Germany. Today, the cost efficiency and high reliability of IPCC make it more appealing than conventional shovel-truck operation, especially in longer life projects with lengthy transportation distances and high production rates (Koehler, 2003). Since 1956, different aspects of this system have been reviewed by many researchers. Barua and Lanergan (1985) developed a computer program that compares the tonnage aspects of various conveyor layouts. They believed that to maximize the economic benefit of lower haulage costs associated with in-pit conveying, one must first minimize the cost of the flatter slope, which is expressed in terms of additional stripping or tied-up ore. Dos Santos and Stanisic (1987) evaluated the design of a high-angle conveyor at the Majdanpek copper mine. Then, Sturgul (1987) applied GPSS (general purpose simulation system) to find the best location of an in-pit crusher."

Jan 1, 2017

By W. Ferreira

Steel is produced in a basic oxygen furnace from hot metal obtained from a blast furnace. A sintered iron ore with good hightemperature properties (strength and permeability) should be used as feed to the blast furnace. The quality of this sintered ore depends on the reactivity of the iron ore used as feed to the sinter plant during the lime assimilation step in the sintering process. The penetration test is the standard method for evaluating the reactivity of iron ore with lime. It is, however, difficult to determine the exact depth of penetration from the standard test. A new test method is proposed that allows automatic evaluation of iron ores in terms of lime assimilation with increasing temperature. A comparison of the coefficients of variation for the new and standard methods for each ore type demonstrates that the results of the new test are more reproducible and more precise than those of the standard method. The test is also less time-consuming and easier to implement.

Jan 1, 2014

By S. Mohanta, I. D. Behera, B. Sahoo

"Run-of-mine coal contains particles of different sizes with different specific gravities, and when this coal is crushed, the amount of material present in a particular density class changes. This alters the washability characteristics of the coal, which affects the quantity, quality, and separation efficiency of the gravity-based coal washing process. The degree of difficulty in washing a typical coal at a particular specific gravity depends mostly on the amount of material occurring within ±0.1 specific gravity range, which is known as ‘near-gravity material’ at that particular specific gravity of separation. In this paper, two numerical indices, namely ‘near-gravity material index’ and ‘index of washability’, are used to quantify the distribution of near-gravity material in different density classes and to evaluate the degree of difficulty involved in the washing process. The results of this preliminary investigation indicate the uniform distribution of near-gravity material in all density classes for finer size fractions, whereas for coarser size fractions more near-gravity material is present in the lower density classes, making the washing process more difficult at a lower specific gravity of separation. IntroductionThe prime objective of coal beneficiation is to remove physical impurities as effectively and economically as possible. The beneficiation process includes various operations on the asmined coal to make it more suitable for enduse application without destroying the physical identity of the coal. There are several coal beneficiation processes, but gravity-based separators remain the most efficient unit operation for removing the undesirable gangue material from run-of-mine coal. These separators exploit the difference in density between coal particles and gangue particles to effect a separation. The feed is introduced into an aqueous suspension of ultrafine magnetite with a particular density. The feed particles with a relative density less than that of the suspension (clean coal) float to the top surface of the suspension, whereas particles denser than the suspension density (gangue material) sink to the bottom. When the difference in density between the particle and the aqueous suspension is large, the separation is easy, but when this difference is less the separation becomes much more difficult since the settling velocity of the particles is very low. This lower settling velocity increases the probability that particles that should report to floats report to sinks, and vice versa, hence increasing the amount of misplaced material. Therefore, the degree of difficulty in beneficiating a typical coal at a particular specific gravity depends on the amount of material occurring within ±0.1 specific gravity range. This is known as ‘neargravity’ material, also termed ‘near-dense’ material, at that particular specific gravity of separation."

Jan 1, 2016

By M. S. Shishvan, J. R. van Duijvenbode, L. M. Cloete, M. W. N. Buxton

Geochemical and mineralogical datasets from Tropicana Gold Mine, Australia, have been used to define ore fingerprints. VNIR/SWIR spectral data were represented by four normalised wavelength regions and were clustered to form spectral classes. Sequentially, these spectral class proportions within a block and collocated XRF data were clustered to form material types (fingerprints). The material types were related to an Equotip-BWi correlation. These correlations can be used to extrapolate a hardness signature and generate a BWi proxy for different blocks. The combined fingerprints and BWi proxy can assist as a tool for enhancing the prediction of comminution behaviour. They can explain specific domain-related hardness variations. For example, one material type could be separated into a softer (~15-18 kWh/t), and harder (>20 kWh/t) material blend. This was accomplished using the commonly overlooked VNIR region at 605 nm. This outcome has significance for blending strategies.

Sep 1, 2021

By C. Norden

The purpose of this article is to show how a representative overall equipment effectiveness (OEE) measurement can be calculated for an underground coal mining bord and pillar batch process. The calculation method, the typical losses (which include planned and unplanned availability losses, coal quality losses, and process rate losses) and the underlying logic are presented and discussed. It is argued that the current traditional underground bord and pillar mining process can currently provide a maximum theoretical OEE of only 49 per cent, while a realistic benchmark target should be in the order of 37 per cent, which relates to an average of 2400 t per shift or 1.2 Mt/a. In addition, it is argued that the current bord and pillar process is the current bottleneck to further improvement past the 2400 t per shift benchmark, and that fundamental process changes will be required to eliminate this bottleneck.

Jan 1, 2012

By D. F. Malan, J. C. Esterhuyse

In this study we explore the use of a novel numerical modelling approach to study the effect of pillar reinforcement on pillar stability. Case studies in the literature indicate that tendons, strapping of the pillars, and shotcrete or thin spray-on liners are commonly used to reinforce pillars. No clear methodology exists to select the type of support or to design the capacity of the support required, however. This has led to ongoing collapses in some mines in spite of heavy support being used to reinforce unstable pillars. A limit equilibrium model with confinement on the edge of the pillar was used to simulate the interaction of the support with the failing pillar. The model correctly predicts that an increase in confinement will lead to a decrease in the extent of pillar failure. As the displacement discontinuity boundary element method allows for the efficient solution of largescale bord-and-pillar layouts, the effect of pillar confinement can now be studied on a mine-wide scale. Accurate calibration of the limit equilibrium model is, however, required before this method can be used for the design of effective pillar support.

May 3, 2023

AFRICAN NATIONAL COMMITTEE ON TUNNELLING In 1970, the South African Institution of Civil Engineers, the South African Institute of Mining and Metallurgy, the South African Institute of Mechanical Engineers, and the Geological Society of South Africa held an international conference on tunnelling, known as Tuncon 1970. SUID-AFRIKAANSE NASIONALE TONNELBOUKOMITEE In 1970 het die Suid-Afrikaanse Instituut van Siviele Ingenieurs, die Suid-Afrikaanse Instituut vir Mynbou en Metallurgie, die Suid-Afrikaanse Instituut van Meganiese Ingenieurs en Die Geologiese Vereniging van Suid-Afrika &apos;n internasionale konferensie in verband met tonnelbou gehou, wat bekend gestaan het as Tonkon 1970.

Jan 5, 1974

By D. Pienaar, C. O’Connor, B. McFadzean

Dithiophosphate (DTP) is classically used as a co-collector in sulphide mineral flotation, where xanthate-DTP mixtures have been shown to synergistically improve mineral flotation performance. Notwithstanding the improved recoveries observed, ethyl DTP does not appear to behave as a conventional collector, since it has been shown that it does not adsorb onto certain sulphide mineral surfaces. Recent investigations have nevertheless indicated that when ethyl DTP does not adsorb onto the sulphide mineral surface, the presence of a non-ionic frother can activate a synergistic increase in pulp phase mineral recovery, and bubble-particle attachment probability. A regular solution approach was used to determine the interactions between thiol collectors and non-ionic frothers at the air-water interface. Attractive interactions were observed between longer alkyl chain DTPs and frothers, whereas the activity of xanthates at the air-water interface were too weak to determine if an interaction was occurring. This is the first time that such interactions have been shown for the DTP-frother system and represents an important contribution to our understanding of the role of DTP reagents in the flotation process. Keywords: Collector-frother interaction, dithiophosphate, surfactant synergism

Jan 1, 2020