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By H. W. J. P. Neomagus, R. K. K. Mbaya, J. R. Bunt, N. T. Leokaoke, N. W. Makgobelele

Silicon Smelters consumes more than 80000 t/a of wood-derived charcoal as carbonaceous reductant in the production of silicon metal. More than 10% of this material is discarded as fines (<6 mm) generated due to abrasion during processing. Charcoal fine residues (<650 μm) and polyvinyl alcohol (PVA) binder were used in this study to produce mechanically strong charcoal briquettes for metallurgical application as carbonaceous reductant. The PVA binder was added in mass percentages of 1, 3, and 5 wt% to the charcoal fines. The compressive strength, abrasion resistance index (ARI), drop shatter resistance (SRI), and water resistance index (WRI) were measured as functions of curing for up to 7 days under atmospheric conditions, and the results compared with metallurgical grade coarse charcoal. The ash content of the produced briquettes was found to be high (6.6–8.0%) compared with the coarse charcoal (1–3%). The 3 and 5 wt% PVA-bound briquettes were found to be the strongest, with compressive strengths of 40 and 115 kg/cm2 respectively, with WRI values of 75 and 73% respectively. The produced briquettes were found to have lower ARI and SRI values compared to the coarse charcoal. Future work should include beneficiation of the fine charcoal discards prior to briquetting, and an increase in binder addition to above 6 wt% to improve the ARI and SRIn.

May 1, 2021

"This paper challenges the traditional notion that mine planners need to plan production so as to incur the lowest mining cost. For a given mine configuration, a mine that increases its mining rate will incur increased mining costs. In an environment in which operations are fixated on cost reduction, a proposal that increases costs will not be readily accepted. Such a proposal requires financial justification—the increase in costs might be recuperated by the additional production. This paper evaluates the net present value (NPV) across a range of copper prices for two underground orebodies located at different depths, using a production rate of 300 kt per quarter and a scenario that introduces additional equipment and costs for 450 kt per quarter. The evaluation was based on the changes of NPV for the orebody located at a shallow depth compared with the orebody at a greater depth. Discrete event simulation combined with mixed integer programming was used for analysis. Unlike traditional sensitivity analysis, this study re-optimizes the mine plan for each commodity price at each production rate. The results show that, for the low mining rate at the final copper price, an NPV of A$1530.64 million is achieved, whereas an NPV of A$1537.59 million is achieved at a higher mining rate. Even though pushing mining rates beyond traditional limits may increase mining costs, this option may be beneficial at certain commodity prices, particularly when prices are elevated. IntroductionA traditional notion among mine planners is that mining rates should be planned in such a way as to reduce production costs. This concept generally relates to the mining rate that results in the highest machine utilization for the selected equipment. Although an existing mine may consider increasing its mining rate by adding equipment to a given mine configuration, this typically increases mining costs. In an environment in which operations are fixated on cost reduction, a proposal to increase costs may be very difficult to argue for. The value of the mine plan is well recognized as varying with the mining rate for a given orebody (Smith, 1997)"

Jan 1, 2017

By M. Hermanus, N. Coulson, N. Pillay

"SynopsisThis paper assesses the effectiveness of the Mine Occupational Health and Safety Leading Practice Adoption System (MOSH) and its potential to improve mine health and safety in South African mines. Developed by the Chamber of Mines, which represents the majority of the country&apos;s large scale mining employers, MOSH was devised to accelerate progress towards achieving health and safety milestones, which were set by tripartite agreement in 2003. The paper documents and builds on the findings of a study conducted by the Centre for Sustainability in Mining and Industry (CSMI) in 2011 that evaluated MOSH strategy, structures, and process of implementation.The study found that MOSH operated across the mining sector, was directed and dominated by experts and, despite best efforts to include other stakeholders, was led by employers. Statutory worker health and safety representatives and structures were not integrated into the complex change process developed by MOSH. The Mine Health and Safety Inspectorate (MHSI) and organized labour were ambivalent about direct involvement in MOSH and preferred regulatory measures to enforce the participation of mines. MOSH interventions were not targeted at mines with a poor health and safety record, and MOSH lacked a baseline from which to track impacts on sector-wide health and safety performance. The leading practices most widely adopted by mines were designed to improve, rather than fundamentally alter, existing practice. Although the depth of engagement with MOSH among stakeholders and on mine sites varied, mining companies, labour representatives, and the Mine Health and Safety Inspectorate (MHSI) saw the programme as significant.IntroductionIn 2011, the Chamber of Mines commissioned the Centre for Sustainability in Mining and Industry (CSMI) to evaluate the progress and effectiveness of the rollout of the Mine Occupational Health and Safety Leading Practice Adoption System (MOSH). CSMI, based in the School of Mining Engineering at the University of the Witwatersrand, is an independent research centre conducting health and safety research relevant to all major stakeholders in the mining sector. The purpose of the evaluation was to identify the factors that enabled or hampered the adoption of new technologies or practices. The timing of the evaluation, two years before the deadline for the 2013 safety and health milestones, allowed time for adjustments to be made. This paper summarizes the results of the evaluation and includes further analysis and discussion of the nature of the programme and the conditions that act to advance it or it hold back."

Jan 1, 2015

By N. Madani, S. Abulkhair

In multivariate geostatistics, it is common to have different types of complexities between variables of interest. In this context, an inequality constraint is an example of complex bivariate relationships. Unfortunately, traditional co-kriging and co-simulation algorithms cannot reproduce this type of bivariate complexity, leading to the overestimation of disturbing elements. This paper proposes a new algorithm based on a hierarchical sequential Gaussian co-simulation framework, integrated with inverse transform sampling, to model inequality constraints between variables. First, the proposed methodology's validity was evaluated by applying it to a real case study from an iron deposit, with an inequality constraint between iron and aluminum oxide. Then the simulated results were compared with a conventional hierarchical co-simulation algorithm to investigate the effect of inverse transform sampling on the quality of the co-simulation. The results showed that the proposed algorithm can reproduce an inequality constraint between variables.

Sep 1, 2021

By J. P. Gouvea

Back-analysis of two mining-induced tremors in 2011 was used to determine a modelling criterion for the entire Klerksdorp Goldfield (i.e., Vaal Reefs region). At that time, most mining operations in the region were owned by AngloGold Ashanti (AGA) and shared many geotechnical similarities. By 2018 AGA had sold off all its Klerksdorp operations to different mining companies, which continued applying the 2011 modelling criterion. As mining activities became more remote, scattered, and isolated, the extraction of isolated blocks of ground and remnants became increasingly necessary. However, the inherited criterion allowed only minimal or no mining activities at these areas. Keeping in mind that safety is the overriding consideration, an innovative numerical modelling approach, which includes the application of peak particle velocity designs, was developed to provide practitioners in the Klerksdorp Goldfield with a suitable criterion to justifiably and safely mine isolated blocks of ground and remnants. This would not always be possible if a conformist approach was used.

Mar 14, 2024

Mr D. Wilson was born in Vereeniging, where he also completed his schooling, matriculating from General Smuts High School in 1977. Mr K.A. van Gessel was born in 1937. His final studies were conducted at Delft University, where he graduated with an M.Sc. degree in mining engineering. Mr J.G. McCaUum was born in 1938 and matriculated at Jeppe Boys&apos; High School, Johannesburg. Mr H. Scot RusseU started his mining career at Randfontein Estates Gold Mining Co. Witwatersrand, Limited, where he served as temporary Geological Assistant, Sampler, Surveyor, Shift Boss, and Mine Overseer. Mr A.M.P. Henderson was born in 1950, in Kiel, West Germany. Mr J.L. Nel matriculated in 1957. This was folowed by his atainment ofthe folowing degrees: B.Sc. (Chemistry, Physics) in 1961, B.Sc. Hons. (Metalurgy) in 1965, M.Sc. (Metalurgy) in 1967, and MDD (student of the year) in 1981

Jan 1, 1989

By M. -H. Fillion

"Geotechnical stability analysis and design in open pit mines requires access to representative geological, structural, hydrogeological, and rock mass models. The quality and quantity of collected geomechanical data used to build such models can have significant implications in the design of safe and economically viable slopes. Data collection is a continuous process throughout the life of a mine. The process starts from exploration to conceptual/preliminary feasibility, feasibility, design and construction, operations (early production/late production), and mine closure.As geomechanical data is often limited at an early stage of a project, the estimated variability of the geotechnical factors may be larger than the actual data variability. This can potentially result in a conservative selection of the pit slope angle. Collecting additional geomechanical data at later stages of the mine development may contribute to reducing the uncertainty associated with the geotechnical properties and a steeper slope angle may be used for the design. This can result in significant economic benefits.This paper investigates the impact of access to new data over time, following a series of geomechanical data collection campaigns, on design decisions. This is illustrated by comparing the resulting slope design, at three stages of the mine project, at a South African open pit mine. To demonstrate this, a series of slope stability analyses were performed using limit equilibrium tools. For each project stage, the slope angle design was modified until a similar probability of failure of 5% or less was obtained. The corresponding factor of safety distributions for each project stage were then compared to evaluate the impact of collecting additional data on the resulting slope design.The results showed that data variability can differ significantly from one project stage to another. Data variability may be greater at more advanced project stages (i.e. a wider distribution of the factor of safety). Furthermore, the results showed that, even for the cases with greater variability, steeper slope angles were possible at more advanced stages for the proposed design requirement (probability of failure < 5%). In this investigation, as more data became available, the uncertainty associated with data variability could be sufficiently reduced to allow the selection of less conservative slope angles. This can be used as a managerial tool in developing data collection strategies and allocating the necessary funds, and can lead to important economic benefits to the mining operation."

Jan 1, 2016

By P. Valicek, M. Farren, R. G. B. Pickering, G. Harrison, P. van Dorssen, W. Joubert, H. J. van Rensburg

Introduction In South African metalliferous mines, stoping operations are largely confined to narrow, tabular orebodies making mechanization extremely difficult. Limited flexibility in terms of stope width has necessitated drilling by means of hand held, pneumatic rock drills at most of our operations. As the largest producer of platinum in the world, we drill in excess of forty million stope blast holes per annum. Cleaning of the broken ore is also done by conventional means—utilizing scrapers and scraper winches. Conventional methods such as these carry high costs in terms of risk exposure as well as labour intensity. The purpose of this paper is to describe the development of a long hole stoping system suitable for the narrow platinum reefs. The new system had to be substantially safer and more cost effective than conventional mining. We selected a site that was conducive to mechanized mining and set out the parameters that we considered would meet our requirements. We then conducted a feasibility study to evaluate the conceptualized long hole stoping system in terms of specific goals. An analysis of the proposed system identified the Key Performance Indicators (KPIs). One of the most critical goals was to be able to create a man-freestope. This would mitigate the safety risk associated with normal stoping operations, as well as the inefficiency cost previously referred to. Other challenges included setting targets for long hole accuracy, stoping width reduction and specific blasting system performance for long hole stoping. We also modelled a critical span that we considered to be inherently safe and to ensure hangingwall stability, throw-blastcleaning and overall system productivity. Anglo Platinum Research and Development and Sandvik Tamrock entered into an agreement for the supply of a specialized long hole-drilling machine. The system was implemented and tested by the mine with the support of the other participants. The results achieved met our expectations. It was demonstrated that with this method of mining it is possible to have a man-free/support-free stope face and to achieve a reduction in shaft head cost per ton whilst reducing dilution. Key performance indicators assessed during the trial •Safety targets •Mining layout •Drill rig •Production targets •Support requirement •Explosives requirement •Services requirement •Ore removal and cleaning requirement •Training requirement Safety targets All persons involved with the project had to adhere to the mine’s safety programme including the zero-tolerance campaign, and were required to report all hazards observed immediately. Mine personnel and Tamrock carried out a full risk assessment on the drill rig before transporting underground. A further risk assessment was carried out when the machine was in position underground. Mining layout See Figure 1 and Figure 2 for a detailed description. Drill rig The drill rig was a combination of a Secoma P-Low LC 10long hole feed and an Axera low profile carrier. The choice of this equipment was made through consultation with Tamrock Sandvik. All relevant design perimeters of the area to be mined were taken into consideration. Production targets

Sep 1, 2001

By C. Steve

ABSTRACT: In this paper, the stability evaluation and bifurcation criteria are presented for the Jinping arch dam structure. The bifurcation models in failure status of the arch dam are derived from the uniqueness of the controlling equation of dam behaviours, and the relative phenomena between positive definition of dam structural tangent stiffness and bifurcation are given in the paper. Then the structural collapse criteria for arch dam are discussed based on the energy mode. A stability criterion of dam based on the connection of yield zones is proposed according to the implementation of elasto-plastic FEM (finite element method) numerical analysis. In accordance with this judging method, nonlinear FEM analysis is used for Jinping arch dam. The failure process of the Jinping dam was simulated and yield zones were found under the overloading condition. Keywords: Stability evaluation, bifurcation, failure process, Jinping arch dam ZUSAMMENFASSUNG: In der Verliegenden Arbeit wird ein integrater geomechanisch Modellversuch fuer Jinping Bogenstaumauer Vorgesstellt. Die Bogenstaumauer liegt aut der komplizerten felsformationen. Der Zerstoerungsmechanismus, die Stabilitaetsfaktoren und die Interaktion zwischen der Bogenstaumauer und den Seiltichen. In der Vorliegenden Arbeit wird eine dreidimensionale finite element analyse fuer die seitlichen. Gestein sauflager mittels dem Bruchmodell Vorgestellt, da fuer Jiniping Bogenstaumauer angewandet wird. Ein Bruch mechanismus model einschliesslich der statistischen parameter der Gesteinfes tigkeit wird entauckeit. RÉSUMÉ: Das ce documènt a été faite une analyse d’elemt limité de trois dimensions dn fondement de construction de roche faisant appel au modéle de l’endommagemnet à fracture, analyse appliquée an fondement du bar ragevoute de Jinping. Ce documènt décrit le test de modéle géomecanique intégral pour le barrage-voûte de Jinping L’ampleur du modéle en question constitute l’une de plus importantes en China. Ils correspondent par faitement anx calculs des résultats.

Jan 1, 2003

By I. Ashmole

The dimension stone industry is not widely known or appreciated within the mainstream mining industry. This is in spite of the fact that the industry has grown at an average of over 7% per annum since 1980, and is estimated to now have a global turnover in excess of US$ 60 Billion per annum. Production technologies, particularly in the developed world where labour costs are high, have developed rapidly over the past two decades. Aligned with this, there has been a significant trend in adoption of a more professional mining approach in the exploration for and operation of dimension stone quarries. In fact, in the experience of the authors, mining a dimension stone deposit successfully if often more challenging from a mining engineering point of view than many conventional surface mines. This paper gives an overview of the industry and then proceeds to discuss best practice in exploration for dimension stone and to outline the latest technologies applied in mining of dimension stone. These technologies are varied and the specific combination of technologies applied in any given quarry is determined by both the physical properties of the material as well as the influence of geology which significantly impacts on the recovery of saleable material and hence the economics of the quarry.

Jan 1, 2008

By D Reddy

Rosh Pinah Zinc Corporation is a zinc-lead mine located in southern of Namibia. It is currently considering the development of a low grade ore body containing about 2.8% zinc. This ore can be utilised by the concentrator if it could be upgraded prior to the ball mill. Preliminary work indicated that the ore may be upgraded in a pre-beneficiating stage by using dense medium separation, X-ray sorting or a combination thereof to a feed grade of minimum 9% zinc. This paper describes the process followed in evaluating the technologies for pre-beneficiation. The results from test work indicated that the low grade ore is amenable to pre-beneficiation by both dense medium separation and X-ray sorting. A comparison between the two technologies showed that dense medium separation marginally outperformed X-ray sorting on grade and recovery. The results for the X-ray sorting were achieved by tuning the calibration curve specifically for the samples tested.

Jan 1, 2008

By J. H. Viljoen, C. H. Graan, N. B. Strydom, J. S. Greyson

M J Martinson (Fellow) There is everywhere a lamentable dearth of quantitative information on occupational factors affecting the health and safety of mineworkers, and as the authors point out in their introduction underground lighting is one of the factors which has received little attention in South Africa prior to this welcome paper. Since the authors are to some extent pioneering, it is perhaps inevitable that the paper should raise more questions than it answers, but nevertheless it provides a provisional yardstick against which the need for, and success of, remedial action may be judged. Before commenting in detail on a few aspects of the paper it may be worthwhile to summarize the statutory provisions relating to underground lighting, contained in Chapter 15 of the Regulations promulgated in terms of the Mines and Works Act 27 of 1956. Chapter 15 is entitled `Lighting, Safety Lamps and Contraband&apos;, but in the context of the present paper only four simple regulations need be noted: firstly, no person shall work or travel in an unilluminated part of a mine unless he carries a light ; secondly, adequate stationary lights shall be provided at stations and certain other localities in shafts, winzes and places used for winding; thirdly, mechanically propelled conveyances shall be provided with an effective bright light shining in the direction of travel; and fourthly, moving machinery to which personnel have access shall be illuminated so as to be clearly visible. In all four regulations the standard of lighting to be provided is wholly subjective, and under these circumstances the wide inter- and intra-mine variations in light intensity disclosed by the study are hardly unexpected. Nevertheless the data presented in the authors&apos; Table I raise three points (1) Is there any evidence that the 19 mines visited during the survey are representative of the 45-50 large gold mines in South Africa, or that the shafts selected are representative of the respective mines? Mine managers would scarcely be human were they not to offer better lit working places for the survey; perhaps the use of a scaling factor (e.g. tonnes or transit mineworkers per shift, kilowatts, etc.) for each working place might have helped to put the figures in perspective. (2) Looking at two of the more numerous - and probably important - of the ten working places included in the survey, i.e. stations and tips, one is struck by the marked. skewness of both intensity distributions (with higher frequencies at the lower intensities) and the abrupt cut-off in both cases just short of pitch darkness. Thus, of the 45 meaned readings recorded at 26 tips no less than 19 fall in the lowest range of 0,1-0,9 1m/ft2, with but one tip in total darkness; of 46 readings on 26 stations, 7 fall in the lowest range and none is in total dark-&apos;1 ness. Although a light with an intensity in the range 0,1-0,9 1m/ft2 may satisfy the regulations, it is certainly arguable that in some situations it may be safer to have no light at all. (3) In view of the skewness referred to above, one wonders how meaningful are the `total average light intensities for (the gold mining) industry&apos; quoted in Table I, which the authors compare with illumina¬tion levels found in other countries. Looking at tips again, one notes that 34 out of 45 meaned intensity readings fall below the `total average&apos; for tips, and only 10 exceed it. From this it may be concluded that a few well-lit working places included in a `total average&apos; may conceal a plethora in Stygian gloom. In their discussion the authors make fleeting reference to the effect of underground lighting on productivity and accident rates, without however quoting any South African experience in this regard. Intuitively it seems highly likely that better lighting (in the widest sense) would improve productivity and reduce accidents, and of course it is in these areas that the mining industry might seek economic justification for providing better lighting. It is all the more regrettable therefore that these factors were not studied simultaneously with the present investigation, especially at the Psychological Division of the Human Sciences Laboratory has recently published a study on underground accidents in the gold mining industry without making any reference whatsoever to lighting as a causal factor. In the event neither of the two studies enables us to make an economic appreciation of the situation, and this seems to be a classic example of a situation where some dovetailing of research effort would have greatly increased the useful return. On the effect of lighting on accidents, it may be noted that in 1939 Mickel presented some very interesting statistics to the Mine Managers Association of the Transvaal2, on the improvement in accident rates which followed the random issue of electric cap lamps to native mineworkers in the Hercules Section of E.R.P.M. the previous year. Based on a 12 month trial period in which 250 000 shifts were worked with carbide lamps and (contemporaneously) 307 000 shifts with cap lamps, the incidence of major accidents (at least one day in hospital among men using carbide lamps was 1,15 accidents per 1000 shifts, while the incidence of minor accidents (requiring dressing only) was 11,52 per 1 000 shifts. The corresponding figures for men using cap lamps were 0,79 and 7,31 accidents per 1 000 shifts respectively, representing a reduction of over 30 per cent in both major and minor accidents. While these figures are obviously dated - only men using cap lamps were issued with hard hats, for instance - they do indicate a need for similar research today.

Jan 1, 1971

The Institute has set important objectives for 1988 to make the Institute still more purposeful. Precious metals The Second International Symposium on Mining, Extraction, and Processing of Precious Metals will be held in Las Vegas, Nevada, USA, during the 118th TMSAIME Annual Meeting in March 1989. The First International Symposium on Precious Metals.

Jan 1, 1987

By C. F. Vos, J. D. Miller, J. C. Davidtz

Small replacements of C12-trithiocarbonate (C12-TTC) for dithiocarbonate (DTC)/dithiophosphate(DTP) in a reagent suite for the flotation of platinum group metals (PGM) have significantly improved the rejection of gangue. Coupled to this is an improvement in kinetics, grade and recovery of PGMs. In a pilot-plant trial, after a 5 mole per cent replacement of the current DTC/DTP collector combination by C12-TTC, large opaque bubbles were replaced by small clear ones. In addition, a similar flotation response was realized with one-third of the current depressant dosage. During a preliminary plant trial on underground UG2 ore, the initial rougher concentrates were characterized by higher PGM concentrate grades and this was in part due to improved chromite rejection. The concentrate produced from the TTC trial also exhibited a much finer particle size distribution. The overall circuit performance, which included secondary regrind and flotation, showed an increase in concentrate grade of approximately 44% while the tailings grade remained unchanged

Jan 1, 2006

A diverse selection of rockbolt designs and resin anchors are available for use in underground mines. Research in recent years at the UNSW Mining Research Centre led to the construction of a rockbolt pull-testing facility. This facility has subsequently been upgraded, commissioned and initial test work has been completed to verify the pull-test process. A test program has been completed with the objective to understand the load transfer mechanism and improve the general performance of rockbolts. This paper describes the results of this research. Une sélection diverse de dessins du rockbolt et ancres de la résine est disponible pour usage dans les mines de métro. Faites des recherches ces dernières années à l&apos;UNSW que le Centre de la Recherche Minier a mené à la construction d&apos;une facilité de l&apos;essai de la traction du rockbolt. Cette facilité a été améliorée par la suite, le travail de l&apos;épreuve en commission et initial a été complété pour vérifier le processus de la traction-épreuve. Un programme de l&apos;épreuve a été complété avec l&apos;objectif comprendre le mécanisme du transfert de la charge et améliorer la performance générale de rockbolts. Ce papier décrit les résultats de cette recherche. Eine verschiedenartige Auswahl an rockbolt-Designs und Harzankern ist dafür verfügbar, benutzen Sie in unterirdischen Bergwerken. Forschung in letzten Jahren beim UNSW-Bergbau Forschung Zentrum führte Ruckerprobung-Einrichtung zur Konstruktion eines rockbolt. Diese Einrichtung ist anschließend befördert worden, ist beauftragt worden und wird abgezeichnet, prüfen Sie, die Arbeit ist vervollständigt worden, um den Ruckprüfung-Prozeß zu verifizieren. Ein Prüfungsprogramm ist mit dem Objektiv vervollständigt worden, den Lastübertragungsmechanismus zu verstehen und die allgemeine Aufführung von rockbolts zu verbessern. Dieses Papier beschreibt die Ergebnisse dieser Forschung.

Jan 1, 2003

By B. D. Paphane

Since the commissioning of the Tati Hydrometallurgical Demonstration Plant (HDP), which produces Ni and Cu cathodes as well as Co salt from sulphide concentrate employing Activox® technology for leaching, no kinetic data has been generated from the plant under normal operating conditions, from start-up to steady state. Therefore, this study seeks to determine the rate of reactions in the leaching circuit with a view to obtaining kinetic data for the system under the current operating conditions. Within the five-compartment autoclave operating at temperatures of 105°C to 110°C and 1100 kPa pressure, the extraction of Ni, Cu, and Co ions into solution is achieved through the oxidation of the sulphide concentrate using approximately 50 kg H2SO4 per ton, 10 kg NaCl per ton, and 1.21 kg O2 per kilogram of S2-. The autoclave reactions showed complex kinetics, with the reactions proceeding via some intermediate(s) before the formation of products. The results suggest that the reactions were dependent on the size of the metal ion of interest and the concentrations of the metal ions in the Tati concentrate as well as in chloride leaching. The order of the leaching rate constants was Fe2+ > Cu2+ > Ni2+ > Co2+ in all the compartments studied. The rates of leaching were also found to decrease as the concentrate moves from compartment 1 to compartment 4. For compartments 1 to 3, leaching was observed to be preceded by dilution of the repulp liquor (copper raffinate). The leaching rate for compartment 5 was slightly lower than that of compartment 1 but higher than the rates of compartments 2, 3, and 4 for all the metals studied. Keywords kinetics, hydrometallurgical plant, Activox

Jun 1, 2013

By D. Marcotte, R. V. Dutaut

"In mineral resource evaluations, geostatistical methods known as global change of support allow prediction of the theoretical histogram and gradetonnage curves prior to interpolations or simulations of grades. Two methods commonly used by professionals to guide the choice of interpolation parameters and assess results are the discrete Gaussian model (DGM) and the indirect lognormal correction (IndLog). These models rely upon an estimate of the dispersion variance of the blocks, which is derived by numerical integration of the variogram model over a discretized block. Due to difficulties in obtaining well-formed traditional experimental variograms (especially in the presence of outliers and limited clustered data), many professionals prefer to use ‘normalized’ variograms such as correlogram (non-ergodic variogram), pairwise-relative, or variogram of the normal score transform. A series of simulations with different grade distributions and variogram models are used to assess the performances and robustness of the various variogram estimators with respect to the DGM and IndLog global change of support. Our results show that the traditional variogram, the correlogram, and normal score variogram have better performances, compared to pairwise, for both DGM and IndLog. Moreover, DGM provided better results than IndLog for the grade distributions that are not strictly lognormal. These findings provide valuable guides for geostatistics practitioners. IntroductionWhen reporting resource estimates,practitioners are required to follow generalstandards described in the NI43-101 (CIMguidelines), JORC, or SAMREC codes. One cangenerally recognize three main parts directlyrelated to the block model:•,Pre-processing: exploratory data analysis (EDA), domaining, capping, compositing, declustering, variography•,Processing: block size, neighbourhood, interpolation types (linear or nonlinear)•,Post-processing: classification, reporting.Pre-processing aims to simplify and strengthen the processing step. Ideally, domaining, capping. and compositing is aimed at defining a single homogenous population. During the processing phase, a common practice is to use a block size corresponding to the planned selective mining unit (SMU). In most precious metal deposits (gold, notably), the SMU size is frequently smaller than the recommended half data spacing (Journel and Huijbregts 1978), resulting in high estimation variance and either a high degree of smoothing if the regression slope is managed or a high degree of conditional bias if it is not. Most actual resource estimates are made using block sizes between one-quarter to one-sixth of the average data spacing, sometimes even smaller. In addition to the block size, the interpolation choices for the neighbourhood selection and interpolator (e.g. inverse distance (ID) or ordinary kriging) are often based on the Qualified Person’s (the QP) experience and some basic validation plots that often do not consider the conditional bias (e.g. trends or SWATHs plots, which are designed to compare two sets of population using a one-dimensional graph)."

Jan 1, 2018

By G. J. van Tonder

The use of Molecular Recognition Technology (MRT), a highly selective separation technology using specifically designed ligands (such as macrocycles), has been demonstrated to be an effective treatment process for removal of several ions that are commonly found in many hydrometallurgical circuits as well as in environ-mental treatment operations. Ions removed by MRT include base, heavy and transition metals, precious metals, halides, and alkali/alkaline earths. A MRT pilot plant run was conducted at Zincor to demonstrate the cobalt/nickel purification of zinc electrolyte, and to recover saleable pure cobalt/nickel sulphate solution. The MRT process holds various advantages, which include the elimination of a catalyst such as arsenic trioxide, conventionally used in cobalt and nickel cementation. SuperLig® 138 can be used to remove cobalt and nickel from120?170 g/l Zn-solutions at pH 1?2. The Zincor MRT application demonstrated its capability of concentrating large volumes o copper barren impure solution at low cobalt and nickel concentration levels in the feed, to small solution volumes at high cobalt and nickel concentrations. Around 99.5% of the cobalt was removed from solution with cobalt impurity levels in feed varying between 10 mg/l and 26 mg/l.

Jan 1, 2002

By G. Konak

In-pit crusher location selection has had an important role recently, since increasing diesel prices. One way of reducing the haulage costs is to shorten the truck haulage distance by bringing the truck dump point into the pit. This study covers the work done on the optimum crusher location definition for a company that produces aggregate near Izmir, a city of Turkey located in the west. This paper will discuss the effects of pit geometry and mine access requirements on optimum crusher location selection that are mainly based on the establishment of minimum haulage distance. The method will be introduced first and then the computer algorithm developed for this study will be discussed later in the paper. Keywords: in-pit crushing, optimum location selection, optimizing haulage, computer aided design.

Jan 1, 2007

By A. Haines, O. Varaud, I. Baidoo

Abosso Goldfields Limited (AGL) is in the process of developing an expansion of their Damang Open Pit gold mine in Ghana. AGL acquired the Damang mine in January 2002, with SRK participating in the geotechnical investigation and design for the project since 2001. A deep resource drilling program, completed during 2004, intersected an extension of the orebody below the then planned final pit floor. A feasibility study has been completed to determine the economic extraction of this deep ore with an expanded pit shell. The Damang mine is in a stockwork sulphide deposit at the culmination of the Damang anticline. The deposit occurs as hydrothermal mineralisation associated with dominantly east dipping thrusts and sub-horizontal quartz veins. Primary gold mineralisation also occurs in the conglomerates of the Tarkwaian Formation. The geological units consist of steep easterly dipping dolerite, sandstone, siltstone and phyllite beds, with the weathering profile comprising saprolites overlying a relatively thin joint oxidized horizon. It is planned to deepen the current open pit to a final depth of up to 270 m, primarily by the development of a push-back of the eastern and western walls. The edge of the Eastern Tailings Storage Facility (ETSF) is currently located within 80 to 100 m of the crest of the existing pit. The geotechnical design focused on the optimization of the bench face angle and finally recommended a ?Steep Slope Strategy?, involving 24m high near vertical benches. Limiting bench stack heights of 96m at a maximum of 66° were derived from the influence of major structures and the variation in rock mass conditions. The overall slope angles in each pit sector were developed around the requirement for haulage ramps. One of the main design challenges was to optimise the standoff distance between the final pit crest and the ETSF. This process had to consider maximum access to the resource without undue influence on or from the nearby tailings dam. A planned increase to the height of the dam wall was an additional factor in the design. Dewatering of the pit walls was also a major consideration in this tropical environment.

Jan 1, 2006