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By W. Newcomen, G. Dick

"Effective management and mitigation of pit slope instability in open pit mines begins with a comprehensive slope monitoring programme. The ability to differentiate between non-critical pit wall movements due to rebound or relaxation of the excavated slopes and movements that may be indicative of slope failure is important for maintaining a safe working environment and maximizing production. Slope failure prediction methods using velocity, acceleration, and strain criteria have been introduced and put into practice over the past few decades. Forty-eight slope failures with surface monitoring data are presented and assessed using the strain-based failure prediction approach, with consideration of the quality of the rock mass and the potential failure mechanism. The results indicate that the strain-based approach can be used to provide general guidance regarding strain thresholds for pit walls for a variety of failure modes in diverse geological environments. The advantages and potential drawbacks of the strain-based and other slope failure prediction methods are discussed. The importance of implementing a pit slope monitoring and performance evaluation system early in mine development is also emphasized. IntroductionThe successful prediction of the Manefay Slide pit slope failure at Bingham Canyon at 9:30 am on 10 April 2013 (Engineering and Mining Journal, 2014) highlighted the importance of having a comprehensive slope monitoring system that provides an accurate prediction of the time of failure of an open pit slope. Using most of the geotechnical monitoring tools available, the time of the failure was predicted within hours and equipment and personnel were evacuated from the pit well in advance of the failure. From a time to slope failure perspective, the monitoring system in place was a success.There are two general groups interested in monitoring pit walls, with slightly different objectives in mind when the monitoring system is being set up. There are those at the mine site that have to implement the pit slope design, monitoring protocols, and the action response plans, and there are those designing the slopes (often consultants) who regularly visit the mine but may not be responsible for the day-to-day activities. was a success.There are two general groups interested in monitoring pit walls, with slightly different objectives in mind when the monitoring system is being set up. There are those at the mine site that have to implement the pit slope design, monitoring protocols, and the action response plans, and there are those designing the slopes (often consultants) who regularly visit the mine but may not be responsible for the day-to-day activities."

Jan 1, 2016

By Z. Leng

"After detonation of the explosive, the final blast-induced damage area around the blast-hole falls into three categories, namely crushed zone, fractured zone, and elastic deformation zone. This paper presents a modified model to calculate the size of crushed zone around a blast-hole in drill-and-blast, with the hoop compressive stress and cavity expansion effect taken into account. The material in the crushed zone was assumed to be a granular medium without cohesion, but with internal friction. On this basis, the formula for crushing zone radius was derived. The proposed approach was verified with tests reported in the literature and the simulated results from SPH-FEM coupled models. A full statistical analysis was also carried out, and the predicted values were found to be in better agreement with the test results compared with other models. A sensitivity analysis of the modified model showed that the size of the blast-induced damage area is mainly affected by the following factors: rock mass properties, in situ stress, and borehole pressure. In particular, the roles of borehole pressure and in situ stress are discussed. Finally, suggestions are made on how the size of the crushing zones can be decreased. IntroductionOver the past decades, drill-and-blast has become the most commonly used technology in rock excavation. It is well known that in rock mass fragmentation with explosives, the annular rocks around the blast-hole are converted into fines. The formation of these fines consumes a significant part of the energy of the detonation, which in general is ignored in the determination of the efficiency of detonation (Glatolenkov and Ivanov, 1992; Furtney et al., 2012). Many studies show that only 20–30% of the total explosive energy is effectively used in fragmenting the rock, and up to 50% of the energy generated by conventional charges is wasted in overcrushing of the crushed zone and the inner part of the fractured zone (Ouchterlony et al., 2004; Sanchidrian et al., 2007). How to control the crushed zone to enhance the effective utilization of explosive energy, reducing the unit explosive consumption and the engineering cost, is therefore of great significance."

Jan 1, 2016

By L. Lorig

"Two types of extreme event, earthquakes and rainfall, potentially affect open pit slope stability. In the case of earthquakes, there are rather well-developed analysis procedures and acceptability criteria. The analysis procedures relate mainly to selection of the dynamic loading either through design earthquakes and/or pseudo-static seismic coefficients. The acceptance criteria are typically expressed in terms of a minimum safety factor in pseudo-static analyses. The acceptance criteria are often promulgated by governments despite the fact that no open pit slope has ever been adversely affected by an earthquake. This paper explains why open pit slopes are seemingly more resistant to dynamic loads than natural landforms, which can experience catastrophic landslides.Extreme rainfall events are much more likely to cause open pit slope problems than earthquakes. Two types of problem are common – slope erosion and slope instability. Slope erosion is often mitigated by appropriate surface water controls. Slope instability due to elevated transient water pressures is more difficult to mitigate. Analysis procedures and acceptability criteria are rare. This paper will discuss the mechanisms for rainfall-induced slope instability, as well as analysis methods. Examples will be discussed and analysis methods are proposed. IntroductionTwo types of extreme event, earthquakes and rainfall, potentially affect open pit slope stability. In the case of earthquakes, there are rather well-developed analysis procedures and acceptability criteria. The analysis procedures relate mainly to selection of the dynamic loading either through design earthquakes and/or pseudo-static seismic coefficients. The acceptance criteria are typically expressed in terms of a minimum safety factor in pseudostatic analyses. The acceptance criteria are often promulgated by governments (e.g., Peru, Chile) despite the fact that no open pit slope has ever been adversely affected by an earthquake. This paper explains why open pit slopes are seemingly more resistant to dynamic loads than natural landforms, which can experience catastrophic landslides.Extreme rainfall events are much more likely to cause open pit slope problems when compared to earthquakes. Two types of problems are common – slope erosion and slope instability. Slope erosion is often mitigated by appropriate surface water controls. Slope instability due to elevated transient water pressures is more difficult to mitigate. Analysis procedures and acceptability criteria are rare, but do exist. In Peru, for example, slopes are required by law to be designed for a 100-year storm."

Jan 1, 2016

"Sampling error and bias, especially the Increment Preparation Error (IPE), are introduced when the tool allocated to do the job fails to extract a representative sample. This is the case with chip sampling; the tool can only extract haphazard shapes of loose or fractured material and not always from within the demarcated sample area. Increment Extraction Error (IEE) and Increment Delimitation Error (IDE) could be severe, but are found to be relatively unbiased. The uncontrolled action of the sampler in discarding excess sample material after sample collection has a significant impact on the error and bias introduced with each and every extraction. A new type of bias, referred to as the ‘waste discard bias’, arises when samplers select what portion of the excess sample material to discard at the sample site. Material visually identified as waste is discarded in preference to mineralized broken ore if there is an excess of sample material. This results in a considerable error in the analytical results. The bias is proportional to the reef-waste ratio in the demarcated sample. IntroductionTwo types of ‘sampling’ are distinguished, namely sampling of particulate materials (broken ores) and sampling of solid, unbroken in situ mineralized reef. This study considers only chip- or face-sampling of in situ, goldbearing, carbon-rich conglomerates typical of the Upper Witwatersrand Supergroup and examines where components of sampling error and bias, clearly specified in the Theory of Sampling, may arise during the practice of ‘face sampling’. Many millions of face samples have been extracted from Witwatersrand-type mines using a hammer and chisel, and detailed descriptions (standard operating procedures) of how the samples should be taken have been developed by many mining companies. There is, however, no taxonomy or nomenclature for managing or reporting the different types of errors associated with chip sampling. The Theory of Sampling compiled by Pierre Gy in the early 1950s provides, comprehensive nomenclature of sampling errors that is applied to particulate sampling of ore in a lot, a stockpile, or a moving stream on a conveyor belt. The taxonomy and characteristics of sampling errors, grouped according to the sources from which they arise, and identified by Gy (1973, 1979, 1982, 1992, 1995, and 1998) are shown in Figure 1. This classification of errors is used extensively in regard to the sampling of particulate materials in the mining and minerals industry, but not in regard to chip sampling. A relatively small mass (hundreds of grams) of representative particulate material for estimating the average grade of a much larger mass (thousands of tons – the ‘lot’ in the Theory of Sampling) is extracted by the chipping or cutting of a representative rock ‘sample’ from the stope face."

Jan 1, 2016

"Pressure leaching is an option for copper recovery from chalcopyrite. Leaching takes place at high temperatures and pressures in the presence of an oxidizing agent. This work shows how the performance of pressure leaching circuits may be improved by optimizing the configuration of autoclaves and the heat removal method. The objective was achieved by creating an autoclave model that combines mass balances, energy balances, and population balances. A base case circuit consisting of a single autoclave was chosen, and this base case was compared with options for increasing capacity by adding more autoclaves to the circuit. These capacity increase options include circuits in which the additional autoclave is added in parallel, in series, and in series with thickening between the two autoclaves. The copper extraction, productivity, and cooling requirements for these options are compared. The series circuit with interstage thickening has the highest extraction and productivity; however, the cooling duty in the first autoclave is high. In addition to comparing circuit options, three options for heat removal were investigated: cooling coils, quench water, and flash recycling. The flash recycle option results in the highest copper extraction, and the quench water option the lowest copper extraction. IntroductionBackgroundCopper sulphide ores and concentrates containing chalcopyrite are typically hard to leach under atmospheric conditions. Smelting of these ores may also be a challenge, due to the inability of smelting processes to handle low-grade ores, as well as impurities such as arsenic and antimony. Alternative hydrometallurgical routes to treating chalcopyrite are therefore required. One option is pressure leaching, which is the focus of this work.In the pressure leaching process, sulphide ores and concentrates are leached in autoclaves at temperatures of 200 to 220°C and a pressure of about 3000 kPa (Schlesinger et al., 2011). An oxidizing agent, usually oxygen, is required to oxidize sulphide ions to sulphate ions. The oxidation of sulphur to sulphate is exothermic; consequently, a method of heat removal is necessary in order to maintain the temperature in the autoclave."

Jan 1, 2016

By J. N. Van der Merwe

"The probability of survival of coal pillars is a more rational indicator of stability than the safety factor. Over time, pillars undergo reduction in size by progressive scaling of the sidewalls. The probability of survival is therefore not constant but reduces over time. The rate of scaling can be quantified, and the reduction of the probability of survival as a function of time can therefore also be quantified. The paper describes the link between the factor of safety, probability of survival, and reduction of the probability of survival. Using the safety factor, probability of survival, and the time at which the probability of survival will reduce to 50% is proposed as a more rational method of evaluating pillar stability than the safety factor alone.IntroductionCoal pillar strength has long been the subject of research in South Africa and several other coal-producing countries. However, it began receiving serious attention only after the Coalbrook disaster in 1960. The most important outcomes of the initial investigations were the groundbreaking research by Salamon and Munro (1967) that resulted in the well-known power formula for pillar strength, and the work by Bieniawski (1969) that resulted in a linear formula for strength. The former approach was based on empirical back-analysis of failed and stable pillar cases, and the latter on in situ strength tests on large specimens.The Salamon and Munro (1967) formula has been widely used in South Africa and even adapted for application to hard rock pillars by researchers such as Hedley and Grant (1972).At the time of the initial work by Salamon and Munro, there were only 27 failed pillar cases suitable for analysis. It was not possible at the time to determine site- or seam-specific strength, and the result was a general formula for use in all seams in all mining districts. Since that time pillars have continued to fail and the database continued to grow – as of 2013, there were 85 cases in the database. Using the additional data, there have been a number of re-evaluations of the strength formula and it has also become possible to differentiate between the different mining seams and areas."

Jan 1, 2016

By M. Pelser, J. Hagemann

"The Anglo American Platinum Rustenburg Base Metals Refiners (RBMR) is the largest producer of nickel cathode in South Africa. It treats slow-cooled matte to recover a platinum group metal-containing magnetic fraction for downstream processing and a nonmagnetic fraction to produce nickel and copper cathode, cobalt sulphate, and sodium sulphate. The refinery was recently expanded from 21 to 33 kt/a nickel. During the ramp-up phase, the sulphur removal section was found to be a constraint to throughput. Nickel hydroxide is precipitated from nickel spent electrolyte by addition of sodium hydroxide, and filtration of the nickel precipitate allows sulphur to be removed from the circuit as aqueous sodium sulphate in the filtrate. The filtration fluxes were found to be limiting the circuit’s capacity. An obvious solution would have been to install additional filtration capacity at significant capital cost and long lead times. RBMR chose rather to first initiate a collaborative project with Anglo American Technical Solutions to better understand the impact of process conditions on filtration characteristics and to evaluate whether the filtration flux could be increased by improving the quality of the nickel precipitate.A deeper understanding of nickel hydroxide precipitation chemistry and kinetics in the context of high nickel and sodium hydroxide concentrations was developed, guided by precipitation theory and confirmed by laboratory tests and plant observations. It was found that the filtration rate could be significantly improved by modifying the precipitate characteristics. A new precipitation process was engineered and retrofitted into the existing plant. Since start-up of the new precipitation system, the available capacity of the sulphur removal section has more than doubled, and control of the process has improved with the better agitation. This paper describes the theory and results used to design the new agitation system, as well as the impact of the change on capital investment, operability, and process efficiency."

Jan 1, 2016

By L. Xingwana

"The purpose of this study is to understand how ore loss and dilution affect the mine call factor, with the aim of subsequently improving the quality of ore mined and fed to the mill. It was necessary to know the relationship between ore mined and delivered for processing in order for the mine to operate efficiently. A review of the literature suggested the mine-to-mill integration approach as a standard intervention to address declining mine call factors. A case study was undertaken at a gold mine in the Free State on the influence of the fragment size distribution on ‘free gold’, the effect of ore loss and dilution on the mill efficiency, and time tracking of ore movement from source to the plant. The study revealed that fragmentation, underground accumulation of ore, and dilution have a significant influence on the mine call factor and mine output. It was difficult to quantify the total ore loss due to the time lag between mining and milling, which is extremely variable, when estimating the mine process inventory. However, it was possible to estimate the total dilution with some degree of precision by analysing the daily blast and head grades. It is therefore recommended that mines accept that some gold loss and dilution will always occur during blasting and transport of broken ore, and this loss must be minimized.IntroductionThe mining of high gold grades is obviously very beneficial for the economic viability of an operation, but optimum recovery of the product is required to maximize profits (Chapman and Bindoff, 2010, p. 13). Although integrating these two requirements is very difficult, due to the various activities occurring between mining and processing, it is indispensable. The main driver of this integration is the impact that different ore types and characteristics have on head grade. According to Isokangas et al. (2012, p. 871), mining and processing operations involve a series of well-defined sequential stages that are complementary in an efficiently operating mine, with the performance of one operation affecting that of another. An attempt to optimize each function separately without considering the whole extraction process will not result in the optimum efficiency of the entire mine. Several researchers have developed a proven methodology for revealing these inefficiencies (Dance et al., 2006, p. 1)."

Jan 1, 2016

By M. H. Dehghanpoor, M. Zivdar, M. Torabi

"A hydrometallurgical route was developed to extract metals such as gold, silver, copper, and selenium from anode slime from Sarcheshmeh Copper Complex in Iran. For the extraction of copper by leaching, the anode slime was heated and stirred with 4 M sulphuric acid in the presence of oxygen. Optimal conditions for recovery of copper were temperature 80°C, mixing time 5 hours, and oxygen flow rate 2 L/min, resulting in a copper recovery of 94.47%. Silver and selenium were dissolved by leaching anode slime from which the copper had been removed with nitric acid. For gold extraction, anode slime was heated with aqua regia at 90°, and the gold then extracted with 2-ethyl-hexanol at a concentration of 0.5 M and phase ratio of 4. The extract was then stripped with soda solution at pH=10.5 and a phase ratio of 5. Finally, gold was recovered by precipitation with oxalic acid at a ratio of mass to volume 10 g/L. Total gold recovery was 80%. IntroductionDuring the electrorefining of copper, some metals are accumulated in the byproduct anode slime, which is an important source for recycling and recovery of these metals (Petkova, 1994). In this paper, a hydrometallurgical route is described for the recovery of precious metals from copper anode slime of Sarcheshmeh Copper Complex, which is located in southeastern Iran .Copper anode slime is black or gray powder with a particle size less than 200 mesh. Approximately 2 to 20 kg of anode slime is produced per ton of copper cathode. The slime is characterized by higher amount of Ag, Se, Pb, and Cu compared to other metals and a very low Au content (Khaleghi and Ghader, 2013).The copper anode slimes are composed of 11.69% selenium, 3.93% copper, 5.26% silver, 0.13% gold, and 45% barite, plus lead, tellurium, antimony, sulphur, and silica with small amounts of nickel, iron, zinc, and bismuth. Gold is one of the most important noble metals due to its wide applications in industry and economic importance, so recovery of gold from slime is more appealing than that of other metals. According to the World Gold Council (2013), the demand for gold has increased during the last decade. This increase is due not only to the increasing demands of the jewellery market, but also to the increasing demand for gold in industrial as well as medical applications (Ranjbar et al., 2014)."

Jan 1, 2016

By C. Chongo, F. X. Paquot, C. Ngulube

"Over the last six years, the Kansanshi copper-gold mine has continued to process significant quantities of complex transitional mixed ores. Initially, the ore treated came predominantly from accumulated stockpiles. Subsequently, the feed source shifted primarily to freshly mined ore that generally showed less tarnish with better flotation kinetics, but with the downside of lower concentrate grades. A number of changes were made to improve the efficiency of the rougher circuit and quality of concentrate produced. These changes were as a result of initial mineralogical studies and laboratory flotation test work that were discussed at length in a previous paper1. Salient among the recommendations adopted as a result of previous work were increased circuit capacity and a modified configuration to increase the residence time and the number of sulphidization stages. These modifications were completed and commissioned in 2011. This, however, resulted in significantly low rougher concentrate grades and a below-specification final concentrate product. This was due primarily to higher mass pulls from the extended rougher bank with multiple concentrate discharge streams, and which necessitated an enlarged cleaning capability. The extra cleaning capacity came on stream in 2014. Of note was the development of a recovery model, based on the strong correlation between the total copper and acid-soluble copper ratio, as a tool for monitoring plant performance. IntroductionThe Kansanshi mine, the largest copper mine in Africa, is 80% owned by Kansanshi Mining plc, a subsidiary of First Quantum Minerals. The remaining 20% is owned by the state through the parastatal Zambia Consolidated Copper Mines Ltd (ZCCM). The mine is located approximately 10 km north of the town of Solwezi and 180 km to the northwest of the Copperbelt town of Chingola. The mine has been in commercial operation since 2005."

Jan 1, 2016

By V. van der Merwe, C. W. Siyasiya

"A petrochemical company experiences premature thermal fatigue failure of the casting moulds used in catalyst production. The aim of the project was to find an alternative alloy that would outperform the current low-alloy cast steel used for the moulds. Based on their thermo-fatigue properties, 3CR12 ferritic stainless steel and H11 tool steel were chosen for testing and comparison with the currently used BS3100 B7 cast steel. Samples of each material were subjected to temperature cycling in a Gleeble 1500TM thermo-mechanical processing simulator, followed by surface analyses. The main parameters derived from the test work were the total true strain, the hot strength of the materials, and the number of cycles to failure. Additionally, the coefficient of thermal expansion for each material was measured using a Bähr dilatometer. H11 tool steel yielded the best performance by way of having the fewest surface cracks, the lowest total true strain per cycle, the most cycles to failure, the highest hot strength, and the lowest coefficient of thermal expansion.IntroductionThe Catalyst Manufacturing Plant (CMP) is a dusty environment with temperatures higher than ambient. Molten mill scale (magnetite, Fe3O4) is tapped from the electric arc furnace into the casting moulds at temperatures close to 1600°C. After 10 seconds the magnetite is quenched with a water spray for 50 seconds to a minimum temperature of 400°C. Sometimes there is residual water in the pans from the cooling fog process (water sprays) in the empty mould. In a typical plant, the factory produces 40–44 batches a day containing between 150–300 moulds, depending on the time required for the specific batch (Buchholz, 2014). The typical batch in the CMP usually weighs about 6 t. The thickness of the tap, speed of the casting belt, and tilting angle all affect the number of moulds per batch.One of the main problems faced during the molten mill scale casting is the premature failure of the casting moulds. Cracks initiate on the surface of the mould, which is in contact with the hot catalyst, and propagate into the interior of the mould. Cracks are generally concentrated in the middle area of the mould where the catalyst makes first contact with the mould, and which experiences the highest temperature fluctuations. This suggests that the cracks are caused by thermal fatigue during the casting process."

Jan 1, 2016

By D. Wines

"Slope stability analyses have traditionally been undertaken in two dimensions assuming plane strain conditions. Although three-dimensional analysis techniques are widely available and used routinely for open pits, two-dimensional analyses are still more common due to the relative ease of model construction and the relatively rapid simulation times.Two-dimensional analyses will often produce different results to three-dimensional analyses for the same slope. It is generally thought that two-dimensional analyses will produce more conservative results. The main reason for the differences is the ability of three-dimensional analyses to account for the three-dimensional nature of the various model inputs, including the slope geometry, the distribution of soil and rock mass domains, the orientation of geological structures with respect to the excavation face, the orientation of the in situ stresses, and the distribution of pore pressure. In some cases, for a long, straight slope in basic geological conditions, two-dimensional analyses can provide a reasonable representation of the problem. However, in many cases, the inability of a two-dimensional analysis to represent the true three-dimensional nature of the problem will lead to unrealistic analysis results.This paper discusses the reasons, as detailed above, for the differences in two-dimensional and three-dimensional analysis results. Work by others is summarized, and additional numerical analyses are performed to provide an improved understanding of the effects of slope geometry, structural orientations, and in situ stresses on predicted stability. Case studies are presented for both stable and unstable slopes, and the behaviour of these slopes is related to the three-dimensional nature of the slope geometry and geology.The previous work, the new analyses performed here, and the case studies show that it is often important to provide a realistic representation of the slope in three dimensions in order to obtain reasonable stability analysis results. This is particularly true for hard rock environments where structurally controlled failure mechanisms are most likely. The paper also highlights the fact that back-analysed properties obtained from one analysis technique are not necessarily applicable to forward analyses using another technique."

Jan 1, 2016

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

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

Jan 1, 2016

By Z. L. Xue, J. L. Li, X. Y. Wang

"In low-carbon energy-efficient basic oxygen furnace (BOF) steelmaking processes, limestone partly or completely replaces the active lime. The effects of limestone calcination temperature (1200–1500°C) and time (5–15 minutes) on lime reactivity and CaO grain size were investigated. The reactivity was evaluated by titration with hydrochloric acid, and the CaO grain size was analysed using scanning electron microscopy. The results revealed that for calcination temperatures higher than 1300°C, the reactivity reached a maximum and then decreased . The higher the temperature, the earlier the peak of reactivity appears. The CaO grains grow with increasing temperature and time, which leads to the decrease of reactivity. Notably, the effects of temperature on CaO grain size and reactivity are more marked than that of time. To obtain active lime calcined at ultra-high temperature by flash heating, the calcination conditions should be 1300–1400°C for 10–15 minutes, or 1400–1500°C for 8–10 minutes. IntroductionLime is one of the most important slag-making materials for dephosphorization and desulphurization in the basic oxygen furnace (BOF) steelmaking process. The reactivity of lime is a measure of its capability to take part in slag-making. Lime having a reactivity of more than 300.0 mL (4 N HCl), determined by hydrochloric acid titration is generally referred to as active lime (Rusjan et al., 2007; Zeman, 2008). At present, the active lime used for BOF processes comes primarily from a kiln or shaft kiln, where limestone is preheated at 700– 900°C, and then calcined at 1050–1200°C (Drenhaus et al., 2010; Oates, 2007)."

Jan 1, 2016

By S. A. Hosseini, O. Asghari

"The quantification of mineral resources and evaluation of process performance in mining operations at Gole Gohar iron ore deposit requires a precise model of the spatial variability of three variables (Fe, P, and S) which must be determined. According to statistical analysis there are complex multivariate relationships between these variables such as stoichiometric constraints, nonlinearity, and heteroskedasticity. Due to the impact of these complexities in decision-making, they should be reproduced in geostatistical models. First of all, in order to maintain the compositional and stoichiometric constraints, additive log-ratio (alr) transformation has been applied. In the next step cosimulation, using stepwise conditional transformation (SCT) and sequential Gaussian simulation (SGS) has been used to simulate multivariate data. Through statistical and geostatistical validations it is shown that the algorithms were able to reproduce complex relationships between variables, both locally and globally. IntroductionThe key steps in mining projects are the quantification of mineral resources, definition of mining reserves, and production scheduling. They rely on the construction of a block model that is used to represent basically the spatial distribution of ore grades (Montoya et al., 2012). The determination of grades and tonnages affects risk assessment and economic evaluation of mining projects. Evaluation of process performance in mining operations requires geostatistical modelling of many related variables (Barnett and Deutsch, 2012). Iron ore quality is characterized by multiple variables: not only the iron grade but also the contaminants that interfere in the subsequent steel manufacturing processes. Consequently, the spatial variability of multiple variables must be determined. Key variables are frequently correlated, and such correlations must be honoured during estimation and simulation. Data from iron ore deposits constitutes compositional data; furthermore, relationships the between assay data are often heteroskedastic."

Jan 1, 2016

By U. W. O. L Vogler

"The behaviour of rocks depends on their material properties, and therefore it is important that these properties are determined by relevant and reliable laboratory testing methods. The ISRM Suggested Methods are very important documents in this regard, and a draft suggested method exists for the determination of the ’complete stress-strain curve for intact rock in uniaxial compression’. The method specifies the required geometry of specimens and the required testing procedures. The research described in this paper involved the determination of complete stress-strain curves for two materials, one exhibiting the Class I characteristic, and the other the Class II characteristic. To obtain a better understanding of the deformation behaviours in the post-peak region, tests were carried out on specimens of both material types with different length-to-diameter ratios. The results showed that the geometries of the test specimens influence the stressstrain curves in the post-peak region. The implication is, therefore, that the numerical values obtained for the slopes of the post-peak graphs are not material constants, but are geometry-dependent. Explanations for this behaviour are given. These findings have important implications regarding the use of post-peak data, for example for blast fragmentation considerations, and for evaluation of rockbursting potential. The research findings are based on a limited amount of laboratory testing, and should be confirmed by further testing. IntroductionThe most common approach for studying the behaviour of rock is by compression testing of cylindrical specimens in the laboratory. Rocks respond to forces by deforming and eventually breaking. The strength and the deformation characteristics are derived from the stressstrain curves produced. Standard shapes and sizes for the test specimens and standard loading rates are recommended in the Suggested Methods of the International Society for Rock Mechanics (ISRM) and the American Society for Testing and Materials (ASTM) standards. In spite of these standards, there remain certain aspects of testing which require further attention. The research described in this paper deals with deformation and failure behaviour of rock materials using stiff, servocontrolled testing equipment. The particular objective was to test the premise that the behaviour of cylindrical rock specimens in the post-peak region of the stress-strain curves does not reflect only a material characteristic, but is influenced by the height-to-diameter ratio (the geometry) of the rock specimen. An additional objective was to investigate the effect of the stiffness of the test equipment on the outcomes of tests."

Jan 1, 2016

By T. Tholana, P. N. Neingo

"Mining companies globally are currently facing severe economic and financial challenges. In addition to global challenges, the South African mining industry has to face other operational challenges that are unique to the country and which threaten the survival and competitiveness of the industry. Profit margins are being squeezed by rising costs and decreasing commodity prices, while labour productivity is greatly affected by intermittent labour unrest. This paper analyses how the South African gold mining industry has performed pre-, during, and post the global financial crisis of 2008. The competitiveness of the industry in terms of labour productivity and industry cost curve position is analysed for the period 2006–2013 to assess the impacts of both the global financial crisis and labour unrest. An analysis of the South African gold mining industry is presented at company as well as mine level. Productivity measure in this paper is limited to labour productivity, in line with limited reporting on productivity. All the data analysed was obtained from the public domain. IntroductionSouth Africa dominated the world as the number one gold producer until 2009 when China took that position, and today South Africa ranks fifth after China, Australia, Russia, and the USA. According to Statistics South Africa (2009), South Africa still had about 30 years of production in the gold sector, a forecast that fluctuates based on the day’s modifying factors. Although South Africa has a comparative advantage in terms of minerals endowment, there are challenges faced by the mining industry that hinder translation of the comparative advantage into a competitive advantage. The industry continually seeks to address these challenges to remain competitive in global markets while addressing national and community needs.Neingo and Cawood (2014) argued that stakeholders in mining are continually concerned with mine management’s ability to translate use inputs effectively into quality outputs. They further suggested close monitoring of effective utilization of people, materials, and money to obtain maximum production of commodities under given conditions. Productivity is measured in various ways, including unit cost, output per employee, and output per unit capital equipment"

Jan 1, 2016

By K. Wu, L. Li, D. W. Zhou, J. W. Yu

The presence of a fractured water-conducting zone (FWCZ) is of significant importance for the safety of underground mining under water bodies. When a coal mining area is covered with a thick layer of alluvial soil, the height, width, and distribution pattern of the FWCZ in the area shows unique features. The internal mechanism by which thick alluvial soil affects the FWCZ is still unknown. Using the Huainan coal mining area (HCMA) in China as a case study, we investigated the impact of thick alluvial soil on the height, width, and distribution pattern of the FWCZ through numerical simulation using the distinct element method, theoretical derivation, and data validation. The results indicate that a thick layer of alluvial soil inhibits the FWCZ height and increases the FWCZ lateral width. When the soil-to-rock ratio is greater than a certain value (the numerical result in this study is 15:18), the FWCZ height and width are affected by the thick alluvium significantly, and the thick alluvial soil alters the shape of the FWCZ from tall and thin to short and wide. The load exerted by the alluvial soil transfers downward to influence the FWCZ; however, this process is obstructed by the hard strata in the bedrock. Therefore, the vertical and lateral change trends show a similar step pattern. Our results can provide a basis for accurate calculation of the FWCZ height in mining areas with thick alluvial soil, to mitigate mine water hazards and increase the safety of underground mining in coal mines covered with a thick alluvium layer. Moreover, reducing the FWCZ height can help extend the upper extraction limit in coal mines covered with a thick alluvium layer, which reduces the thickness of the impermeable coal pillar and increases coal production.

Jan 1, 2016

By T. Moyo

"The oxidative leaching of chalcopyrite in ammoniacal solutions has been evaluated using electro-analytical techniques and controlled bulk leaching studies. The anodic dissolution process has been established to be a sevenelectron transfer process under nitrogen in ammonia–ammonium sulphate solutions and ammonia–ammonium carbonate solutions, which suggests that the sulphur is oxidized to thiosulphate and the copper and iron released as Cu+ and Fe2+ in these systems. The deportment of Fe2+ and S2O3 2- is affected by choice of the ammonium salt used in the leaching process. In the perchlorate salt, only five electrons are transferred, supporting formation of different sulphur species as proposed for the sulphate and carbonate salts. Scanning electron microscopy and energydispersive spectroscopic analysis of the mineral surface after leaching indicate presence of an iron–sulphur surface layer completely free of copper in the sulphate system, an iron-rich surface layer in the perchlorate solutions, and absence of surface layer build-up in carbonate solutions. XRD analysis of a bulk leach residue from leaching in ammoniaammonium sulphate solutions showed the surface layer to be mostly amorphous (90%). The crystalline content (10%) is composed of 95% polymorphs of anhydrous iron oxide hydroxide FeO(OH). Choice of ammonium salt and the hydrodynamic environment of leaching have been shown to influence the presence or absence of the surface product, as well as its nature. IntroductionHydrometallurgical processes for copper extraction provide a viable route to recover the metal from mixed and low-grade ores, as well as overcome the environmental challenges faced by the traditional pyrometallurgical processes. Chalcopyrite is not only the most abundant of the copper sulphides, but also the most stable, making it recalcitrant to hydrometallurgical processes. Hence, hydrometallurgical processing of chalcopyrite continues to be an attractive area of research due to the vaguely understood surface chemistry of the mineral in different aqueous media. Different routes for hydrometallurgical treatment of chalcopyrite can be followed. These include thermal treatment prior to leaching, direct leaching and direct electrochemical leaching (Venkatachalam, 1991). Direct leaching of chalcopyrite can be carried out in various solution systems, as reviewed by Roman and Benner (1973) and Venkatachalam (1991). Ammoniacal solutions are attractive and effective lixiviants that form stable amine complexes with some base metal cations while rejecting iron. Leaching of chalcopyrite in ammoniacal solutions in the presence of an oxidant is possible due to the stabilization of copper (I) and copper(II) by ammonia at elevated pH levels. In oxygenated ammonia solutions, it has been suggested that chalcopyrite dissolves according to Equation [1] (Beckstead and Miller., 1977a):"

Jan 1, 2016

By C. Birch

"The purpose of this study was to establish the impact of discount rates on cut-off grades for narrow tabular gold deposits as characterized by the goldfields of the Witwatersrand Basin in South Africa. There are various methods available for determining the cut-off grade, from simple breakeven calculations to sophisticated software packages that consider a variety of inputs to optimize the cut-off grade. For this study a simple financial model was created in Microsoft Excel® that links the ore flow, block listing, and the cash flow. This allows the cut-off grade to be optimized considering how the cost of capital and chosen discount rate affect the cash flow.The discounted cash flow (DCF) and resultant net present value (NPV) are a widely used valuation method for production properties according to the South African Code for the Reporting of Mineral Asset Valuation (SAMVAL Code). The financial model in this study utilizes the Solver function, as well as simple Microsoft Excel spreadsheet formulae to optimize the NPV. Solver was chosen as it is a standard feature in Excel and thus no additional software costs are incurred beyond the basic Microsoft Office suite. For the purpose of this study, just narrow tabular gold deposits of the Witwatersrand Basin were considered. An example of a typical ore block listing, as well as the costing figures, was obtained from an operating gold mine. The results obtained from the study financial model were compared to the current cut-off grades obtained from the mine using their proprietary optimizer program, and were found to be comparable. The methodology utilized for this study thus appears valid.The cut-off grade was optimized considering the cash flow, which includes the variable mineral resource royalty tax, the variable gold income tax, as well as the discount rate. By comparing the resultant NPVs using discount rates of zero, 9%, and 12%, the impact of the discount rate on cut-off grades, resultant life of mine, and average mining grades (AMG) could be compared for the example ore block listing. The South African Code for Reporting of Exploration Results, Mineral Resources and Mineral Reserves (SAMREC Code) does not require operations to state if a discount rate was considered in determining mineable reserves. Mining companies approach the question of applying discount rates in determining their cut-off grades, and thus mine planning strategy, in different ways. Due to the negative impact that optimizing purely on NPV has on the life of mine, some choose to optimize only on profit. Other companies have reverted to utilizing pay limits as their primary grade planning strategy. The calculated cut-off grade is then considered to be another planning indicator rather than a hard determination of mineable reserves."

Jan 1, 2016