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The AMIRA Code of Practice and Guidelines for Metal Accounting was developed in response to demand from a number of sponsor companies, including one major platinum producer. Since its publication in 2007, the Code has become widely accepted in the mining industry world-wide as a guide for best practice in metal accounting. With the changes in the structure of the platinum mining industry in South Africa over the last 10 ? 15 years, in particular, the starting up of a number of small platinum producers who must have their concentrate smelted and the contained valuable metals refined by third parties, the need for metal accounting systems that are accurate and have an acceptable level of precision has become even more critical to the success of a platinum-producing operation. The current state of metal accounting in the industry is discussed and possible areas for improved accounting and associated metallurgical efficiencies are identified.

Jan 1, 2012

By A. Chuma, K. J. Nyembwe, H. Simonsen, J. H. Potgieter

Different concentration techniques were evaluated for preconcentration of a mineral ore at a coarser size to avoid energy and resource wastage. Specifically, the aim was to reduce milling costs and energy required in the beneficiation of tin. In this study, cassiterite (0.17% Sn) was the mineral of interest in a pegmatite ore body associated with quartz (SiO₂ > 60%) and alumina (Al₂O3 > 20%). Three concentration techniques, namely dense media, shaking table, and flotation, coupled with characterization analysis, were used to assess the concentration response. The results confirmed that particle size and mineral liberation impact the separation process. High recovery and grade were obtained with gravity concentration methods (dense media and shaking table) for coarser (+300 to +212 μm) and intermediate (+150 to +53 μm) particle sizes. Lower recovery and grade were identified for much finer sizes (−53 to −38 μm). Flotation produced a high-grade product at a relatively low recovery and appeared to be only applicable to finer grains. Separation efficiency based on Schulz’s equation measured a segregation performance of 74.8% for dense medium separation and 60.7% for the shaking table for the coarse and intermediate particle sizes. Flotation only achieved a separation of 30%–40%. The results suggest that use of dense media separation as a rough preconcentration method prior to further grinding, and the utilization of a more advanced concentration technique for mineral recovery and upgrade, constitute a successful approach to improve process economics.

Apr 14, 2024

By M. A. Belokrys, F. O. Milovich, D. D. Aksenova, V. A. Ignatkina, V. A. Bocharov, A. A. Kayumov

Copper remains one of the most sought-after metals in the industry. Tennantite (Cu12As4S13) and tetrahedrite (Cu12Sb4S13) that occur in pyrite copper, copper-zinc and polymetallic ores are an additional reserve for increasing technological parameters of flotation. Pyrite mass (85% and more) has similar flotation properties to other sulfide minerals, and imperfection of reagent regime is a common problem that occurs during quality copper concentrate processing. Due to tennantite's and chalcopyrite's differences in flotation properties with butyl xanthate at pH > 10, tennantite often reports to the flotation tails. Hight alkalinity of the medium increases the oxidation level of sulfide minerals, thus increasing the concentration of S2O3 2-, SO3 2- SO4 2-. At processing facilities in the Russian Federation, the concentration of thiosulfate ions in pulp is in the range of 0.5 to 1.5 g/l and higher, which leads to a decrease in flotation activity of the tennantite. The effect of butyl xanthate (as standard), thionocarbamate, dithiophosphate and M-TF (a mixture of dithiophosphate and thionocarbamate) on monomineral (tennantite and pyrite) flotation in the presence of sodium thiosulfate was studied. The variable parameters of the "thiosulfate-collector" flotation complex were thiosulfate concentration (100-900); pH (8-12,5). The scientific methods included contact angle for hydration assessment; TEM for surface studies; UV spectroscopy and iodometric titration for solubility control. Comprehensive studies have established a connection between the concentration of thiosulfate ions and tennantite flotation selectivity. Flotation activity of tennantite is higher at pH 8 with a thiosulfate concentration under 300 mg/l. Pyrite remains active with butyl xanthate under the same conditions. TEM images of the surface films of the collectors on pyrite were acquired at pH 8. The MTF provides the biggest difference in adsorption, wettability and flotation activity between tennantite and pyrite at pH 8 and thiosulfate concentration under 300 mg/l, in comparison with butyl xanthate. The results of the research will be applied to improve selective reagent regimes of tennantite minerals flotation. Keywords: Tennantite, pyrite, chalcopyrite, sphalerite, flotation, sulfhydryl collectors, thiosulfate, adsorption, mechanism, reagent regime

Jan 1, 2020

By B. D. H. Knights

In November 2009 Mintek installed the FloatStar? advanced flotation stabilization system on Vale's Cauê iron ore beneficiation plant in Brazil. In September 2010 the system was upgraded to include flotation optimization. The circuit consisted of two parallel cleaner circuits producing final concentrate. The tailings from the cleaner section passed through a scavenger circuit. Large, well-instrumented feed sumps played an important role in circuit stability, presenting a good opportunity for advanced control. Previously the plant was controlled using distributed control system (DCS) level control only. The flotation circuit processes roughly 30kt/day. The FloatStar system provided advanced control of the flotation circuit as well as of the sumps feeding the circuit. The system underwent a lengthy trial to assess the benefit that it provided. Two data sets were analysed, showing performance under advanced stabilization and optimization respectively. Several validation and consistency criteria were employed to ensure the quality of the analysis. Iron recovery and iron tailings grade were used to measure system performance. The analysis showed that the system increased recovery by up to 2.7%. In addition, the system decreased the iron tailings grade by between 1.2% (from 23.5% to 22.3%) and 4.3% (from 31.3% to 27.0%) during different test campaigns. The analysis also showed that the entire FloatStar stabilization system was active for 72.7% of the time from first activation to the end of the first data set (approximately 200 days). From the analysis it was concluded that under similar conditions for 'ON' and 'OFF' tests, the system provides a clear benefit. It was also found that over the longer term, the system continues to provide a benefit. This finding suggests that the results were not simply due to a short-term advantage. The size of the data set, as well as the magnitude of the recovery improvements, lends considerable confidence to these results. Therefore, it can be concluded that the FloatStar control system provides substantial benefit to operations at this site.

Jan 1, 2012

By N. B. Devi

The extraction of Cu(II) from HCl media using Cyanex 923 in kerosene was investigated. The effect of shaking time, aqueous phase acid concentration, extractant concentration, chloride ion concentration, various metal ions (Mn2+, Mg2+, Fe2+, and Ni2+), and temperature on the extraction were studied. Using 0.1 M Cyanex 923 the extraction of Cu(II) was 6.7 per cent from 3 M HCl, which increased up to 78.5 per cent with 6.5 M HCl and then decreased due to acid extraction by the extractant. The maximum extraction (97.5 per cent) was obtained with 1 M Cyanex 923 from 5 M HCl solution. A mathematical model has been developed to describe solvent extraction equllibria over wide ranges of concentration of extractant, acid, and chloride ions. The composition of the extracted species has been predicted on the basis of slope analysis results and extraction equilibrium modelling. The negative values of enthalpy change and entropy change confirm the exothermic formation of the extracted complex.

Jan 1, 2012

By N. Sacks, N. B. S. Magagula, I. Botef

"Low-pressure cold gas dynamic spraying was used to deposit WC-4wt%Ni and WC-4wt%Ni-1wt%Mo coatings onto mild steel. Dense coatings with very low porosities were produced. No decarburization occurred during deposition and no deleterious phases were formed. The coatings were subjected to standardized material characterization tests, as well as slurry abrasion testing to assess their wear behaviour. The wear tests were conducted in synthetic minewater-silica slurries, while distilled watersilica slurries were used as a control. The hardness of the coatings, 512HV0.3 and 458HV0.3 for WC-4wt%Ni and WC-4wt%Ni-1wt%Mo respectively, are comparable to those achieved using high-temperature coating processes. The abrasion wear rates for both coatings were less than 5 mg/min and 10 mg/min in the distilled water-silica and synthetic mine water-silica slurries respectively. The approximately 50% increase in wear rate in the synthetic minewater slurry is attributed to a synergistic corrosive wear mechanism. The predominant wear mechanisms were identified as binder smearing and delamination, with carbide grain fracture and pull-out. IntroductionSlurry abrasion is a common phenomenon in mining operations, particularly in pipelines transporting, for example, sand-water slurries. This type of wear can cause pipeline fracture, shortened operational lifetimes, and impede pipeline efficiency. To reduce or limit abrasion, thermally sprayed tungsten-carbide-based coatings are often used as pipe linings (Kanchan et al., 2010). This is due to their attractive mechanical properties, which include high wear resistance, stiffness, hardness, compressive strength, and tensile strength. These properties are provided by the coating microstructure and composition, which is a blend of hard, wear-resistant tungsten carbide (WC) particles embedded in a fracture-tough metallic binder phase (Brookes, 1997). Cobalt (Co) is the most commonly used binder due to its excellent wetting of the WC phase and moderate mechanical properties; however, its application is limited to non-corrosive environments as it has a low corrosion and oxidation resistance (Andren, 2001). Nickel (Ni) may also be used as a binder, and has been shown to have superior oxidation and corrosion resistance, but in some applications may not have the required mechanical properties (Kim et al., 2006)."

Jan 1, 2016

By J. U. Burger

The Elizabeth Bay aeolian deposit in Namibia comprises a sedimentary package of diamondiferous grits that have been deposited, reworked and concentrated to various degrees by dominant southerly winds. Diamond concentration in these deposits is highly variable and often unpredictable. The deposit is marginal from an economic perspective and therefore requires selective mining which in turn requires correct sampling and accurate estimation. Three methods of sampling have been used at Elizabeth Bay for mineral resource estimation: (i) pit samples utilizing an excavator to produce a sample 1 to 3 metres deep with a 10 to 20 m2 footprint, (ii)GB50 grab sampling with a 3.34 m2 footprint and (iii) a BG36 bucket auger drill with a 5 m2 footprint. Pit sampling, being easy to execute, has been the most widely used despite a limited penetration ability and, at times, questionable integrity. The GB50 grab sampling tool was tested at Elizabeth Bay and although it had better penetration ability and sample integrity, it struggled with cemented horizons. No sample optimization was conducted prior to the GB50 grab sampling. The latest phase of sampling was conducted using a BG36 bucket auger drill which is capable of penetrating thick overburden, penetrate cemented horizons and provides an acceptable sample support size. Geostatistical simulation methods were applied to establish the optimum sampling parameters required to estimate a mineral resource at an indicated level of confidence. The results show that a 5 m2 sample support size is adequate and that a 50 m 100 m sample configuration is optimal. Sample data obtained from the three methods confirms that the BG36 bucket auger tool combined with the parameters identified in this exercise produce the best results for diamond resource estimation.

Jan 1, 2009

By I. Nolet

Matte tapping is one of the most critical operations of a matte smelter. Every smelter has developed particular matte tapping practices and tap-hole repair strategies that are often based on the experience of the smelter personnel. As a result, very little information has been published on these critical topics. This paper aims at providing the information gathered from a survey conducted in the PGM and Ni industries on matte tapping practices and tap-hole maintenance strategies for different PGM-Ni matte smelting operations.

Jan 1, 2014

By E. J. Margi

Deep gold exploration such as is used for Witwatersrand gold deposits is both expensive and time-consuming. This paper provides quantitative guidelines as to the economic optimum number of boreholes and deflections within each borehole for a certain reef characterized by its variability pattern and depth below surface in order to maximize the level of confidence for a fixed drilling budget or to minimize the total drilling cost for a given level of confidence. Portions of the Basal, Vaal, and Elsburg Reefs that show a wide range of variability (sample logarithmic variances of 0,63, 1,05, and 2,06 respectively) are examined at depths below surface ranging from 1000 m to 4000 m over a simulated mining property covering 7000 ha (7000 m by 10000 m). It is concluded that four intersections per borehole are optimum in al the cases considered, which cover the majority of current Witwatersrand exploration, with the exception of the highly variable Elsburg Reef at depths of 1000 m and 2000 m, where boreholes with one and two intersections respectively should be preferred.

Jan 1, 1987

By R. Castro, R. Gómez, F. Betancourt, M. Moncada

In block cave mining, rock fragmentation is a key parameter that influences the production level design and mine planning. Fragmentation occurs mainly by natural breakage during the caving process and in the draw column. The breakage that occurs within the column is known as secondary fragmentation. Secondary fragmentation has been successfully described using the block caving comminution model, which replicates the fragmentation mechanics between particles under drawn and vertical loads in a draw column. This model is based on a kinetic and population balance approach, in which non-normalized and normalized assumptions can be used depending on material and comminution system behaviour. In this paper, the non-normalized and normalized approaches are applied and compared to laboratory data to determine which assumption should be used for secondary breakage in block caving. Approaches are compared graphically, with the mean square error and the Fisher test with a false-rejection probability of 0.05. Based on a statistical analysis, the results show that the normalized model can be applied to all the rock types tested.

Nov 1, 2021

By D. F. Malan, P. M. Couto

We propose a layout design for the UG2 Reef where weak geological alteration layers are present. The collapse of the Everest platinum mine in South Africa indicated that these layers substantially weaken the pillars. The popular Hedley and Grant pillar strength formula cannot be used where these alteration layers are present. Underground investigations at Everest mine and numerical modelling of the layout were conducted using the TEXAN code and a limit equilibrium model. Simulations of a collapsed area and an intact area allowed for a preliminary calibration of the model. This was subsequently used to explore modified layouts for these ground conditions. An alternative is to compartmentalize the blocks of ore using barrier pillars. The numerical modelling predicted that the barrier pillars appear to remain stable even in the case of large-scale collapses, provided their width exceeds 25 m. Main access routes into the mine can be protected by a double row of pillars at least 15 m wide to provide a safe travelling way. As a cautionary note, these conclusions are based on the model calibration and this needs to be refined in future. Calibration of the limit equilibrium model remains a challenge owing to the large number of parameters involved.

May 3, 2023

By N. V. Moraka

Mining companies are required by law to improve representation of women on their boards. However, progress in this regard has been slow. Boards of directors play an important role in formulating corporate strategy, risk assessment, and effective governance for sustained financial performance. Although some women have successfully maintained board seats, others have been unable to do so. It is difficult to ascertain whether women are being appointed as tokens to satisfy social pressures, as little is known about the experiences of women board members in mining, which may inform or refute tokenism. In-depth interviews were conducted with 20 women and 16 men across six listed mining companies. Thematic data analysis revealed evidence of tokenism where recruitment to mining boards is compliance-based and is informally driven by influential directors. Adverse boardroom experiences that further confirmed tokenism were reported by women, such as disregarding their contributions, condescending behaviour, and limited influence in decision-making, while other women sought to validate their competence and oppose social exclusion. This study recommends that a strong Board Chair and nominations committee is critical to ensure sustainable recruitment of competent and suitably qualified women; and further foster a culture of inclusivity and valuing gender-diverse boards.

Dec 13, 2023

Book new a Book reviews Economics of the mineral industries. 4th Edition. Society of Mining Engineers (Caler No. D, Litleton, CO 80127, U.S.A.). 660 pp. Reviewer: M.A. Ford Foundry technology, by P.R. Beeley. Durban, Buterworth Scientific, 1982. R124 + GST. Reviewer: A.P. van Wyk Underground Transport Symposium. Johannesburg, The South African Institute of Mining and Metallurgy Special Publication SP 1. 300 pp. R65 (Non-Member), R55 (Member), + GST. Reviewer: C.L. Workman-navies A summary of the Project DevelopmentSymposium as presented by The Australasian Institute of Mining and Metalurgy. $A 25.00. Reviewer: D.J. van Niekerk Geology of uranium deposits. Proceedings of the CIM-SEG Uranium Symposium, September 1981, edited by T. I. I. Sibbald and W. Petruk. Quebec, The Canadian Institute of Mining and Metallurgy, Special Volume 32, 1985.268 pp. Canadian $40 members, $50 non-members (plus $3 for postage and packaging). Reviewer: M.S. Toros Crushing and grinding process handbook. Chichester (England), Wiley, 1986. £49. Reviewer: M.P. Hay O'okiep's 50th birthday The 0'okiep Copper Company Limited, situated at Nababeep s,ome 35 km from Springbok, recently celebrated its fiftieth anniversary.

Jan 1, 1987

Written contribution by J.F. Curtist The paper is an interesting exercise in the application of mathematics to a theoretical situation presumed to oc¬cur underground. However, some of the assumptions on which the calculations are based are unlikely to occur in any actual mining situation. Assumptions made in the paper Objections to these assumptions are noted seriatim. (1) `it is assumed that any loading-rate effect will be small and can be ignored' (p. 3). A loading rate of the order of 50 kN-12 per cent of the maximum load-is customary to prevent the props being dislodged by the blast. (2) `it is unlikely that all the props will be installed with their full stroke available, but for the purpose of this analysis, it is assumed that this is the case' (p. 3). A more reasonable assumption would be that, with the most careful selection of prop length and the most rigid control of stope width, the available stroke length would on average be reduced by the order of 70 mm on installation-20 per cent of full stroke. (3) `each support unit supports a block of 3,3 m2 bound¬ed by vertical fractures with zero cohesion that are parallel and at right-angles to the stope face respec¬tively' (p. 5). Joughin and Jager have shown this to be an unwar¬ranted assumption in respect of the positioning, the direction, and the dip of the fractures present in a stope in rockburst conditions. If, as assumed, there was `zero cohesion' between such blocks, the removal of a hydraulic prop in row 3 supporting such a block would ensure that it fell out before a profile prop could be installed in the same position. (4) `the height of each of these blocks is 3,0 m, the volume of each block is 9,9M3, the mass of each block is 27 225 kg' (p. 5). The concept of individual props being affected by dis¬crete blocks of known and uniform dimensions of relatively low ground velocities (less than or equal to 2,16 m/s) acting in line with the axis of the prop is one that lends itself to neat mathematical calcula¬tions. It does not, however, bear much similarity to the actual conditions that are exposed underground in excavations where rockbursts occur. Ortlepp and co-workers concluded that `the impulse of the shockwave can cause the displacement of the

Jan 1, 1991

By J. K. Pothal, S. Basu, R. Barik, P. K. Senapati, A. Dubey

The aim of this paper is to study the characteristics of limestone particles (size varies from 4 mm to 20 mm) during its transportation from open cast mines through pipes. The influence of four different size fractions of limestone samples (4–6 mm, 6–10 mm, 10–15 mm, and 15–20 mm) on pipe flow characteristics during vertical transport in a 100 mm NB hose pipe was investigated. The flow velocities in the vertically lifting pipe was varied from 0.95 m/s to 2.78 m/s and a rotary feeder of 2 to 4 tonne/hr. capacity was employed to feed the limestone samples to the lifting pipe. The pressure drop data for the different size fractions of limestone samples at different mixture velocities were analysed. The effect of input volume fraction and superficial velocity of limestone samples on hold-up ratio were discussed. It was indicated that there was no substantial increase in pressure drop of slurry samples in the studied range of concentrations (CW = 4.22% - 12.54%), particle size, and low to medium range mixture velocities with respect to water run head loss data. The hold-up for the larger size fractions (15–20 mm) of limestone samples was found to be significant as compared to smaller size fractions, which may be due to terminal settling velocities of individual fractions of the studied samples.

Mar 4, 2025

By L. Chanderman, C. E. Dohm

"Mineral Resource estimation (MRE) and classification are highly dependent on the confidence in the model of the orebody geometry, known as the geological model, within which the estimates are constrained. An understanding of the geology of a deposit is thus fundamental to the Mineral Resource evaluation process. The geological model is a function of the structural and depositional complexity of the geology captured by drilling information, which is limited in the early stages of a project. The information effect has a significant impact in terms of the interpretation of the geology as well as the estimation of grade continuity and the resultant risk associated with the Mineral Resource estimate and classification.Therefore, as new drilling information becomes available, it is necessary to update and refine the geological model to appropriately constrain the resource estimates. The grade information is also updated, including statistics and the spatial correlation characteristics. This refinement is based on geological interpretations derived from the drillhole information that becomes available throughout the life of the mine; orebody knowledge that comes from working experience of the deposit, as well as familiarity with the software package used for modelling, is also important.In updating, the potential to identify new resources in previously unknown areas may be realized. The opposite may also occur where previous interpretations of ore continuity are refuted. Both scenarios could result from a better understanding of the geology of the deposit and thus a more realistic MRE, which reduces geological risk. This paper describes how newly drilled advanced grade-control holes for a gold deposit in Mali were used to refine the 3D geological model for the deposit in an attempt to identify the potential for additional oxide ore using conventional geostatistical evaluation techniques."

Jan 1, 2017

By G. J. McTeer, D. J. Rioux

Teck Metals Ltd operates an integrated zinc-lead complex at Trail, British Columbia, Canada. Since 1997, KIVCETTM flash smelting technology has been utilized for the treatment of zinc plant residues and lead concentrates. Trail Operations’ furnace uses four water-cooled jackets for bullion tapping. This paper provides an overview of KIVCET bullion tapping for the past 16 years, and discusses successes and challenges encountered with respect to operations, training, and monitoring.

Jan 1, 2014

By E. M. Sacris

The semi-autogenous grinding mill is the bottleneck whenever treating hard and finely fragmented ore at Benguet Corporation's Kaline Concentrator. Low milling rates result from very high and fluctuating mill power draw. Attempts to use a mill horsepower control loop to regulate feed rate were unsuccessful because of operational upsets caused by fluctuations in feed rate on downstream operations. The objective of the computer control was to maximize tonnage throughput by stabilizing the power draw at a high but safe level without upsetting subsequent operations. The semi-autogenous mill is operated in closed circuit with a stationary screen. Instrumentation consists of a feed weightometer, a power transmitter to measure mill load, a headwater flowmeter and an ultrasonic level sensor to measure the mill discharge sump level. A 64 k-byte FOX-3 Foxboro computer processes field data from various sensors and manipulates the input control variables. These variables are the fresh feed rate and the headwater flowrate. Feed rate is controlled through a supervisor-type computer program aimed at maximizing and stabilizing power draw. The control scheme was according to how the operators responded during an operator-directed automatic control mode. Mill dilution is similarly controlled. Results achieved reflect a 3% increase in tonnage milled because of the reduction in control variability that ensued.

Jan 1, 1987

By R. Simon, P. Pagé, L. Li, P. Yang

The Mathews-Potvin stability method is widely used in the Canadian mining industry as a starting point to determine the maximum dimensions of mine stopes. However, it cannot be applied to inclined (more frequently encountered) mine stopes without conducting numerical modelling to obtain the stress factor A, defined as a function of the ratio of unconfined compressive strength of intact rock to the induced principal stress on the exposed stope walls. The need to conduct numerical modelling significantly limits the application of the Mathews-Potvin method. In addition, given its empirical nature and main application for preliminary design, it is deemed undesirable to conduct numerical modelling, especially elaborate modelling. Alternatively, theoretical methods can provide a much simpler and quicker way to estimate stresses around stopes and the corresponding stress factors. Over the years, a large number of studies have been conducted to estimate stresses around openings excavated with various crosssections. However, theoretical or graphical solutions remain unavailable for mine stopes that typically consist of horizontal floor and roof, and two parallel inclined walls (hangingwall and footwall). To remedy this situation, a series of numerical simulations is first performed for openings with vertical and inclined walls, including typical stopes commonly encountered in underground mines. A group of empirical solutions is then formulated to estimate the induced principal stresses at the roof centre and mid-height of the stope walls. The validity and predictability of the proposed solution have been verified using additional numerical simulations. The proposed solution can thus be used to calculate stresses and the resultant stress factors A around typical mine stopes with any inclination angle and height to width ratio, under any in-situ stress state, without conducting numerical modelling.

Aug 1, 2021

By Y. Rajpal

The Hedley and Grant pillar strength formula is widely used within the South African mining industry to calculate pillar strengths in bord and pillar mining layouts, and is currently being used at Bathopele Mine. Bathopele is an operation of Anglo American Platinum near Rustenburg, mining the UG2 reef at depths ranging from 40 m to 300 m below surface. The formula includes a parameter K, which is a downrated uniaxial compressive strength value of the rock, used to calculate the strength of a pillar. At Bathopele Mine, for the shallower areas Less than 150 m the value was 33 MPa, but this was increased to 44 MPa due to the fact that rock mass conditions improved and there were no incidence of failure in the shallower areas. Underground observations showed pillars to be stable with no signs of failure, and it was questioned whether the design was not still too conservative, resulting in sub-optimal extraction ratios and consequently a loss of revenue to the mine. A trial mining section was started at the mine where the pillars were reduced from 6 m x 6 m to 5 m x 5 m and the pillar behaviour was monitored for the onset of failure. Numerical modelling was conducted using the TEXAN code to back-analyse pillar stresses in order to find a more realistic value for K. Results from the trial section indicated that the pillar design was in fact too conservative, and the value for K was adjusted to 54 MPa for future designs. This paper details the findings from underground monitoring and numerical modelling results, and includes details for further calibration of the K value in order to further optimize the extraction of the orebody.

Jan 1, 2012