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By M. Tang, X. Tong

"SynopsisSlime coating is one of the most common ways for serpentines to contaminate metallic mineral concentrates during traditional flotation of coarse sulphide particles. This could pose quite a complicated and challenging problem in the case of some types of low-grade and finely disseminated Cu-Ni ores bearing high serpentine contents. This is the case for the copper and nickel sulphides from the Yunnan Mine, China. Previous batch flotation tests of this ore resulted in satisfactory recoveries of 86.92% Cu, 54.92% Ni, and 74.73% Pt+Pd, and concentrate grades of 4.02% Cu, 3.24% Ni, and 76.61 g/t Pt+Pd. However, the MgO content in these concentrates was more than 19%. In the current study, microflotation tests and settling rate tests were introduced to investigate the effects of a combination of strong collectors (a 2:1 weight ratio of butyl xanthate and butyl ammonium dithophosphate) on entrainment of serpentines in metallic mineral concentrate, as well as visual observations of the concentrates in suspension using still photography. All test results indicated the presence of serpentines entrapped in the hydrophobic flocs that resulted from these collectors, even with the use of effective gangue depressants. These strong collectors are used to flocculate the ultra-fine sulphides by forming loose and ‘fluffy’ hydrophobic flocs. However, these hydrophobic flocs may also be able to load or entrap some serpentine slimes into the concentrate, and this entrained serpentine could be harder to remove by using depressants or intensified conditioning than serpentine slime coating on the particle surfaces.IntroductionSerpentine minerals, which have the generalized composition (Mg,Fe)3Si2O5(OH)4, can be easy to crush and grind due to their convoluted and bent layered structures. These serpentines can hinder the enrichment of some metallic minerals and dilute their concentrates by entrainment during flotation. The resulting high pulp viscosities, slime coatings, and high content of dissolved ions adversely affect the flotation recoveries of copper and nickel sulphides, and high levels of MgO entrained in the concentrates lead to heavy penalties from smelters. Since the quality of nickel flotation concentrate depends heavily on its MgO content (generally less than 6–7% for No.1 or 2 grade nickel concentrate according to the standard requirement from National Nonferrous Industry), reducing the entrainment of serpentine during flotation is becoming more and more pressing. Previous research has focused on the collectors and depressants that are commonly used in copper-nickel flotation, as well as their interaction mechanisms."

Jan 1, 2015

By C. Tokoro, T. Kato

The objective of this study is to improve the shrinking core model for particles with a particle size distribution and clarify the activation mechanism of rare earth elements dissolution by mechanochemical activation. To achieve this objective, the method, called the doubly integrated micromodel, was applied for a shrinking core model in this study. This method consists of integrating the batch leaching rate of a single particle over the distribution of both feed particle size and reactor residence time to arrive at the fraction unleached. The usefulness of the shrinking core model improved by doubly integrated micromodel was confirmed by comparing it with a shrinking core model. To apply the shrinking core model improved by doubly integrated micromodel to rare earth elements dissolution from weathered residual rare earth ores by mechanochemical activation, leaching experiments were performed for grinding samples by planetary ball milling. The reaction equation of mechanochemical reaction by planetary ball milling for weathered residual rare earth ores was investigated by x-ray absorption fine structure analysis. As a result of a kinetics study, the rate-determining process of leaching changed from reaction to diffusion reaction control by planetary ball milling. Based on these results, it was revealed that the main mechanism of activated rare earth elements dissolution from weathered residual rare earth ores was a mechanochemical reaction by planetary ball milling. Keywords: Mechanochemical reaction, particle size distribution, shrinking core model

Jan 1, 2020

By D. Krebs, Tero Korhonen, X. Yang, A. Kirpala, N. Heino

At present, over 85% of the world’s rare earth oxides are produced from the bastnäsite ores, found mainly in China. Meanwhile, the ion adsorption ores are the most important sources for heavy rare earth production. As the demand for rare earth materials in the world increases, other forms of rare earth deposits will become new potential resources of rare earth production in the future. In this paper, flotation performance of a steenstrupine ore from the Kvanefjeld deposit for rare earth recovery was studied through laboratory bench testwork. The effects of various factors including grinding size and collector dosage on recovery and selectivity of the selected rare earth elements, La, Ce and Y, were investigated. A rare earth concentrate with the grades of La 2.8 wt%, Ce 5.1 wt% and Y 1.0 wt% at recoveries of 80.5% La, 79.8% Ce and 55.3% Y was obtained in optimum conditions. Further potential to improve the efficiency of REE flotation was discussed. Keywords: Rare earth elements, flotation, steenstrupine, Kvanefjeld deposit

Jan 1, 2020

By A. Drumond, J. F. C. L. Costa, M. G. Lemos, A. L. Rodrigues, F. G. Niquini

The past decade has seen a renewed possibility of using machine learning algorithms to solve a large collection of problems in several fields. Data acquisition for mining operations has increased with the growth in sensor-based technologies, and therefore the amount of information available for mining applications has dramatically increased. Ore sorting equipment is available for separating ore from waste based on differences in physical properties detected by a real-time analyser. The separation efficiency depends on the contrast in these properties. In this study we investigate the application of machine learning models trained using data from the output of a dual-energy X-ray ore sorting apparatus at a gold mine. The particles were first hand-sorted into ore and gangue classes based on their mineralogical composition. Classification models were then used to help decide the balance between the number of true and false positives for ore in the concentrate, with a view to economic parameters, using their receiver operator characteristic (ROC) curves. The results showed AUC (area under the ROC curve) scores of up to 0.85 for the classification models and a maximum reward condition Fpr/Tpr around 0.5/0.9 for a simplified economic model.

Aug 8, 2024

By J. Pelders, F. Magweregwede, S. M. Rupprecht

Increased depths of conventional gold and platinum mines, longer travelling times as working places move further from the shaft, and ancillary activities taking place in the mining cycle reduce the useable time spent at them workface. Current shift lengths of 8 hours 20 minutes may no longer be adequate ton complete required mining activities.The impact of mineworker commuting times on shift work is also not well understood. The aim of this research was to draft a framework to optimize shift cycles in order to maximize productivity, health, safety, and wellness in the South African mining sector. The framework was informed by a review of the literature, case studies at a local gold mine and platinum mine, and a workshop with industry experts. The findings indicate potential advantages of extended shift lengths and continuous operations, such as improved productivity, asset utilization, and employee morale. However, concerns included fatigue and the associated safety hazards, particularly when working longer shifts, night shifts, and numerous consecutive shifts. Job demands, personal factors, and commuting times are further considerations when scheduling shifts. The ideal shift system for the mining sector should holistically consider and balance operational requirements, health and safety, and employees’ and their families’ requirements, and suit the unique needs of each operation. Changes in shift cycles will require a structured change management process, which involves all stakeholders, and could help to enhance the viability of the South African mining sector.

Aug 1, 2021

By D. F. Malan, P. J. Le Roux

Almost no work has been published on the strength of pillars in mines exploiting the LG6/LG6A chromitite bands in the Western Bushveld Complex. The strength of these pillars is unknown, and the hard rock industry still uses the Hedley and Grant formula. Numerical modelling, using inelastic constitutive models, may be of some value in estimating the pillar strength, but this approach is difficult and prone to errors as many assumptions are made. This paper explores the alternative approach of the back-analysis of actual LG6/LG6A mining layouts, using displacement discontinuity codes to simulate the average pillar stresses on a mine-wide scale. From this, and underground observations, a ‘minimum’ pillar strength (K-value in the pillar strength formula) can be estimated. The codes TEXAN and Map3D are both used in the paper to simulate the actual pillar shapes and sizes in an LG6/LG6A mine. The results are encouraging as the two codes produced similar APS values for a complex pillar geometry. An estimated K-value of 77 MPa was determined for the LG6/LG6A pillars in this particular area. This is more or less similar to earlier work done for the UG2 chromitite seam, but higher than the K-value of 39 MPa for a linear formula obtained by other workers for LG6/LG6A pillars. Caution should therefore be exercised before using this higher value for all areas. The effect of element sizes and explicitly simulating the effect of the surface for shallow pillars is illustrated in the manuscript using both codes. The need for actual pillar observations and an iterative design process, which cycles between modelling and observations, is also emphasized.

Nov 27, 2024

By B. van Beek, J. H. Zietsman, A. Lexmond, N. J. Andrew

"In platinum group metals (PGMs) smelting, the slag-to-matte ratios are very high (approximately 4:1 to 9:1). The slag temperature therefore has a strong influence on matte temperature, which is a continual concern due to the low matte melting temperature and its resultant high superheat. Maintaining a stable slag temperature is therefore a critical safety and operational issue. This project aims to improve the understanding of the slag bath temperature drivers so that a more effective slag bath temperature control strategy can be implemented.With the aim of developing a multiphysics model of the furnace, extensive data analysis was performed to develop an in-depth understanding of the process before the start of modelling work. In this work, the effect of fluctuating furnace feed composition on slag temperature was examined. The results are based on equilibrium calculations using FactSage, implemented with ChemApp.A method was developed, using an equilibrium calculation at 25°C, to convert the available chemical assays to approximate mineralogical assays in order to determine the feed enthalpies. Actual 2013 furnace feed data was used and the average specific energy requirement to produce slag and matte at 1600°C was calculated to be 671 kWh/t concentrate. The furnace energy input was then kept constant, and it was found that the slag temperature varies significantly as a function of the feed composition. These results were compared to actual slag temperatures measured on site. It was found that close management of the furnace energy balance is important for maintaining a stable slag temperature."

Jan 1, 2014

Dr Hanekom I would like to thank Dr Hanekom for a very valuable -contribution; it summarizes careful statistical test work which was an omission from the original paper. Minor manipulative details, the application of a finer measuring graticule and the reproducibility of the procedure are described and the author accepts with gratitude this ready-made answer to the first and second points of Dr Finkelstein's comments. Mr Williamson The author is indebted to Mr Williamson for historical background to early experiments with captive bubbles and it is conceded that the original paper described another avenue to an old flotation goal. The field of application is restricted and we would not work with all the material illustrated by Klassen and Mokrousov.1 The particles in Figure 74(a) are far too large and irregular; material in Figures (b) through (d) could be employed for semi-quantitative work in spite of obvious disadvantages; Figure 74(f) represents ideal conditions. We have studied this mineral-collector association in the size range given and results were satisfactory. For larger pick-up values than that illustrated an induction period of more than 0.01 second was required (actually 90 seconds for a maximum bubble load at pH 7.2 with 25 g/t oleic acid!). REFERENCE 1. KLASSEN,V. I. ANDMOKROUSOV,V. A., 'An Introduction to the Theory of Flotation,' Butterworths, London, 128, Figure 74, (1963). Mr Slander and Mr Kooij Researchworkers in a practical mineral development laboratory may well ask what contribution yet another micro-technique can make to the problems of large scale ore flotation. The obvious answer is that there can be little or no direct translation of results obtained with a static system comprising air bubble and pure mineral concentrate to a dynamic multicomponent operation. However, for the rapid study of one mineral or variable some small scale apparatus must be employed. In the Anglo American Research Laboratory, the described semi-quantitative pick-up technique has filled a niche and requirement. For ourselves, the system represents a definite advance over the contact angle approach in that samples of minerals conventionally prepared in laboratory or plant may be used, which is a desired step nearer flotation. We bracket this pick-up apparatus with the Hallimond tube1,2 and evidence is now to hand that pick-up curves may often be related to flotation recoveries obtained from this tube with pure mineral samples and from the Fuerstenau glass cell3 with the total ore at low pulp densities. We would not wish to push the comparison of pick-up results past this stage for there is much published development work relating the Hallimond and Fuerstenau equipment to flotation on a larger scale. At the Anglo American Research Laboratory for example, a test in a Fuerstenau cell is often the final integrated step in laboratory investigations into the TORCO process. As these contributors rightly point out, pick-up tests can offer nothing towards the determination of optimum reagent dosage for large scale flotation. This is the function of bench-type flotation cells with up to 2 kg batches of ore. At this level, collector performance and consumption and physical properties of the froth are more comparable with large scale operation. The pick-up cell, Hallimond tube and Fuerstenau cell are essentially equipment from which the effect of control parameters can be derived. For the assessment of collector behaviour, the pick-up approach is rapid and economic and we offer here a pick-up contribution towards the flotation of Witwatersrand uraninite with anionic sulphonate collectors (Fig. 1). A large crystal of the local mineral was not available for contact angle measurements and owing to its low concentration in Witwatersrand gold-bearing reef, a small sample was laboriously obtained by tabling, superpanning and heavy medium separation. The pick-up values of a sized portion were determined by the described procedure in the presence of nine petroleum sulphonate collectors. Responses fell into two categories; "normal" curves (dotted lines) and a group of collectors which were very active at low concentration but showed depression of pick-up at the higher levels. The similarity in form of curves for collectors 1 and 3 was apparent and reagent 3 was found to be simply a diluent of 1. Otherwise, collector performance could not be related to either sulphonate content or molecular weight of the active constituent. However, reagent 2 gave the best performance in a bench flotation cell with regard to overall uranium distribution but, as predicted, reagent 1 yielded the best grades. Later pick-up tests with a similar class of collector have suggested that depression at high reagent concentrations may be due to "bubble armouring" whereby an hydrophilic air-liquid interface is produced which rejects mineral attachment. Prolonged investigations with the same mineral sample have not been required except for the study of uraninite which lasted for more than a year. Variation in pick-up response could not be detected over this period during which the mineral was stored under distilled water. The only visible change was a small efflorescence of lead sulphate, readily removed and decanted off after stirring with a glass rod in the sample tube. In its main role as a "trouble-shooter", the pick-up method is applied to fresh mineral samples taken from a flotation plant and soon discarded or from a current run in the ore-dressing laboratory. After production of a concentrate, a solvent cycle, as described in the original paper, may be required to remove heavy organics or other contaminants as when a sample has been taken from circuits containing detergents, grease or return flotation solutions. This solvent cycle will not remove chemisorbed collectors and we prefer not to work with minerals from flotation tailings or with pick-up cell discards. It has been thought preferable to answer the contributors' other apprehensions with regard to sample preparation, activators, depressants and the physical properties of reagents in another paper on offer to this Journal. It covers pick-up investigations relating to the flotation of nickeliferous pyrrhotite at two Anglo American Corporation Mines, which approach seemed the correct orientation for a reply to another contributor to this paper.

By J. Zwartendyk, A. Azis

"Every nation must decide how it should exploit its mineral resources so as to contribute most to its national objectives. It is argued that the present value of the future corporate income tax flow derived from mineral exploitation may serve as a useful indicator in moving toward a taxation policy that would contribute optimally toward the greatest public benefit.The future stream of corporate income taxes depends largely on the mineral endowment, consisting of deposits now being mined, known unmined deposits, and unknown deposits.Unknown deposits may be estimated by geological inference. Both known and unknown deposits can be categorized according to grade and size, and divided into economic and uneconomic deposits on the basis of certain prices, costs, rate of return and taxation levels.Computer simulations can be run of discounted cash flow models at various taxation levels to show the impact of taxation changes on future tax revenues."

Jan 1, 2014

By A. E. Landsberg, K. C. Corin, G. Stroebel, J. A. Miller, M. Becker, H. J. J. Gordon

Geometallurgy represents an integrated approach to mine planning and operation whereby the naturally occurring ore variability and its effects on all key downstream activities in the mine life cycle are considered. Process mineralogy is one of the key components of this integrated approach. This paper focuses on understanding some of the mineralogical variability across the Swartberg Cu-Pb-Zn sulfide ore in South Africa and its consequent effects on flotation performance. This process mineralogy project is motivated by the planned changeover and expansion of the Black Mountain Mines concentrator to process solely Swartberg ore compared to the run-of-mine blend currently being treated at the plant. Eight samples spanning the main lithological units of the ore body are described using quantitative mineral-textural data and bulk sulfide batch flotation performance to establish five process mineralogical or geometallurgical units comprising: quartz-pyroxmangite-magnetite, quartz-amphibolemagnetite, quartz-magnetite, garnet-magnetite-quartzite and mica-quartzite ores. Important mineralogical characteristics affecting Cu, Pb and Zn recovery, grade and selectivity specific to the different ore samples are discussed. Ultimately, the use of ore-specific process mineralogical and geometallurgical knowledge allows for better mine planning and plant optimisation. Keywords: Flotation, polymetallic base metal sulfide ore, process mineralogy, ore variability

Jan 1, 2020

By T. Flitton, R. Peattie

"The extreme depths and consequent expense of drilling and sampling the gold-bearing reefs of the Witwatersrand Basin have resulted in limited data being available for estimation of grade ahead of the mining face. There is, however, a wealth of information from mined-out areas of these deposits. This estimation challenge resulted in the development of a unique method of Mineral Resource estimation.AngloGold Ashanti (AGA) utilizes a technique termed macro cokriging (MCK), which allows for the integration of the limited advanced borehole data with the large chip sample data-sets from the previously mined-out areas by adopting a Bayesian geostatistical approach. The MCK process, in short, is the estimation of mixed support size data together with the application of four-parameter distribution models.The gold value estimation for the Carbon Leader Reef (CLR) on the AGA TauTona Mine is used in this paper as a case study of the process and to demonstrate the effectiveness of this technique through production reconciliation. The current method has been proven over the last 20 years and is now an established part of the Mineral Resource evaluation process within AGA.IntroductionOwing to the extreme depths of the goldbearing reefs at the AngloGold Ashanti (AGA) Witwatersrand operations and the prohibitive expense and time involved in drilling boreholes to the required depth, only limited drilling and sampling data is available ahead of the mining face. Whereas traditional estimation techniques are interpolative, within the Witwatersrand the estimation is primarily extrapolative, with the majority of the data being sourced from the mined-out areas. This significant challenge to estimation resulted in the development of a unique method of Mineral Resource estimation and evaluation."

Jan 1, 2016

By M. Mpofu, G. Stenzel, F. Magweregwede, B. Maphalala, T. Kgarume

Surface coal operations in which mining activities are conducted above old underground workings experience hot holes after drilling. In a coal mine, hot holes are defined as shot holes, which after being drilled have an in-hole ambient temperature of 40°C or above or show a temperature increase of 3°C or more during monitoring. Hot holes and other cavities, such as cracks, pose health and safety risks to workers, such as exposure to hot air and high concentrations of noxious gases released from these holes. In addition, workers may be exposed to premature detonation of explosives caused by in-hole temperature increases and chemical reactions. To this end, Coaltech Research Association commissioned a project to develop a Best Practice Guideline for the management of hot holes. This paper is a compilation of the activities conducted in the development of the Best Practice Guideline between 2021 and 2022. The activities included a review of standard operating procedures and hot-hole temperature-measuring and -monitoring devices, and the assessment of current hot-hole procedures at selected mines. The results indicated that management of hot holes requires a focus on pre-emptive risk assessment of mining blocks, identification of hot holes using the correct temperature-measuring devices, and continuous monitoring of hot holes from the time of drilling until just before blasting. Hot-hole management accessories, such as polyvinyl chloride sleeves, were found to be effective in insulating the hot-hole emulsion from the rock mass temperature, thus preventing the potential for premature detonation.

Sep 9, 2024

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 F. S. Hoseinian, D. Deglon, L. leal Filho, M. Safari, T. C. Souza Pinto

This paper investigates the effect of chemical and physical operational parameters on the reverse flotation of Timbopeba itabirite iron ore in a pilot-scale pneumatic flotation cell. Operational parameters include collector and depressant dosages, feed rate, superficial gas velocity and solids concentration. Results show that increasing collector dosage results in an increase in quartz recovery, particularly for the coarser quartz where recovery increases by nearly 30%. Increasing depressant dosage results in a decrease in both quartz recovery and hematite loss, for the coarser hematite particles. Increasing the feed rate has no effect on quartz recovery or hematite loss, suggesting that the pneumatic flotation cell is not residence time constrained. Increasing the superficial gas velocity results in an increase in quartz recovery to an optimum at a Jg of around 0.6 cm/s after which it decreases significantly, particularly for the coarser particles where recovery decreases by as much as 50%. Decreasing the solids concentration does not affect quartz recovery but results in significant decreases in hematite loss. Here, fine and coarse hematite loss decreases from around 25% to 12% and 10% to 2% respectively. Superficial gas velocity and solids concentration are operating parameters which have the largest effect on the flotation performance of the pneumatic flotation cell. Results suggest that this flotation cell should be operated at relatively low superficial gas velocities and solids concentrations. However, these are pilot-scale findings and would need to be tested at an industrial-scale where different operating conditions may be optimal. Keywords: Operational parameters, reverse froth flotation; entrainment; iron ore

Jan 1, 2020

"This study examines the problems faced by open pit mine superintendents who make choices about how to direct their materials, either to the waste dump or to the mill. The paper explores the effects of introducing a 10% sampling error and a 0.9-times to 1.1-times sampling bias on positively skewed distributions for precious and base metals, negatively skewed distributions in the case of bulk commodities, and normal distributions as is the case for coal deposits. Parent distributions for each commodity were created on a 25 x 25 m grid using transformations of gold, iron ore, and coal data-sets, spatially based on a nonconditional Gaussian simulation. Ordinary kriging of grades for the three commodities into a 10 x 10 m grid provided the reference case against which the distributions with the sampling error and sampling bias for the commodities were compared. Imposing cut-off grades on the actual-versus-estimated scatterplots of the three commodities allowed the distributions to be classified into components of waste, dilution, ore, and lost ore. Ordinary kriging of values for each deposit type acted as the reference data-set against which the effects and influence of 10% sampling error and 0.9-times to 1.1-times sampling bias are measured in each deposit type. Indications are that the influence of error and bias is not as significant in gold deposits as it is in iron ore and coal deposits, where the introduction of small amounts of error and bias can severely affect the deposit value.IntroductionThe effects of poor sampling and the financial implications for mining companies, traders in mineral assets, and sellers of metal as dore or commodities is documented in a number of studies, the most notable of which is that by Carasco (2004). He examined the financial impact of poor sampling practices in the Chilean copper industry and found that finacial losses due to poor sampling amounted to hundreds of millions of dollars over the life of a mining operation. Holmes (2004) examined the effects of correct sampling and measurement as the foundations of metallurgical balances, and found that revenues from the sales of large iron ore shipments may be profoundly affected by poor sampling practices. In South Africa the way in which sampling of in situ gold-bearing reefs affects the mine call factor has been an ongoing study since the disparity between the estimation of gold in the reef and the actual gold bullion produced was noted by early investigators such as Beringer (1938), Jackson (1946), Sichel (1947), Harrison (1952), and a number of others."

Jan 1, 2017

By B. Maybee, J. Venter, J. Wesseloo

The selection of open-pit slope angles involves high-value decisions with small changes in slope angle, often representing significant changes in net present value. The existing methods used to define design acceptance criteria have evolved from factor of safety to probability of failure (P[F]) to risk consequence to risk frontiers, which represents current state of the art. Current design acceptance measures are based on references to published tables representing the experience and judgement of their authors, who did not specify the purpose of the measures contained therein. As the purpose of the published design acceptance criteria is not specified (i.e., definite stability, marginal stability, optimization, etc.), such tables cannot be used to achieve optimum slope angles that maximize profitability. This paper develops a design acceptance criterion that maximizes profitability by defining a formulation for optimum risk that balances expected risk and reward. The model developed is titled the Mining Risk Model (MRM), and is applied to open-pit slope angle selection through an equation for optimum probability of downside (P[D]O) that balances the upside or opportunity, the P[F], and the downside impact. This formulation for optimum risk is unique, as many authors have presented objective functions for their risk models that can be optimized, but none of the sources reviewed contained a formulation for optimum risk. The MRM is sufficiently flexible to allow the design performance measure that drives P[F], and hence P[D]O, to be selected based on the intended goal. Furthermore, the essential information that must be known to quantify optimum risk is defined. This allows users to determine what information to collect for optimum risk decisions. A further benefit of the MRM is that slope angle decisions and pit shells can be ranked to select the best option, and a threshold is provided that separates acceptable from unacceptable decisions. Finally, the workflow and information required to determine optimum risk are presented.

Sep 9, 2024

By J-C. Lv, S. Zhao, F-L. Zhang, T. Yan

"To eliminate nozzle clogging during continuous casting of ship plate steel and reduce the calcium addition required, the composition of CaO-MgOAl2O3 inclusions in the liquid phase region must be controlled. Thermodynamic calculations using FactSage show that when the balanced Al content is 0.03 mass %, aCa in the liquid steel is 1–10 ppm, and aMg is 0.01–0.3 ppm; and when the balanced Al content is 0.06 mass %, aCa in the liquid steel is 1–30 ppm, and aMg is 0.05–0.9 ppm. The CaO-Al2O3-CaS inclusions can be controlled in the liquid region below 1600°C at aCa values of 5 ppm, 15 ppm, and 30 ppm, and the balanced Al content should be controlled in the range of 0.001–0.050, 0.005–0.250, and 0.010–0.50 mass %, respectively. Based on theoretical analysis, the refining slag composition should be controlled to 45–53 mass % CaO, 30–37 mass % Al2O3, 6–12 mass % SiO2, and 6–8 mass % MgO. Optimization of the slag system and steel composition can reduce, or even eliminate, the requirement for calcium treatment. The total oxygen content in the slab can be controlled to 15–21 ppm, with typical micro-inclusions ??10 ??m in size, with compositions in the CaO-Al2O3-MgO and CaO-Al2O3-CaS systems. Introduction With the rapid growth in steel output in China and the changes in the macroeconomic environment, competition in the steel industry has become more intense. The technology for economical cleaning of high-grade steel is important for improving product quality and reducing production costs. However, the functions of refining equipment have not been fully taken into consideration, and the combination of refining methods has not been optimized completely. These problems have become a bottleneck, restricting modern steel enterprises from developing an economic clean steel production process. Ship plate steel produced by the converter long process has been selected as the example in this study. The composition of the steel is shown in Table I. This steel is a high-strength low-alloy steel for rolling media and heavy plates, mainly used for ocean-going vessels and other steel structures. The carbon content in the steel is far from the peritectic point, and the crack sensitivity of the slab surface is weak, so impurities and gas content are strictly limited. Therefore, the process route BOF (basic oxygen furnace) – LF (ladle furnace) – CC (continuous casting) needs to be studied."

Jan 1, 2018

By V. A. Leão, V. S. Coimbra, V. A. A. Oliveira, R. F. M. de Souza, G. M. de Lima

A willemite concentrate consisting mainly of willemite (55.3%) and dolomite (15.6%) was reduced in a tube furnace with charcoal as reductant. The effects of temperature and varying reductant additions on zinc extraction were studied. A zinc recovery of 86% was achieved with 20% reductant in the charge at 1100°C. During reduction, the calcium oxide generated through calcination of the dolomite promoted decomposition of the willemite, thus lowering the temperature at which reduction took place. The presence of CaO improved the zinc recovery at all temperatures studied. The best zinc recovery (93%) was obtained in the presence of 5% CaO, 20% charcoal reductant, and at a temperature of 1100°C. The results show that the willemite concentrate studied has great potential to be used as a raw material for the Waelz process.

Nov 6, 2023

By D. F. Malan, W. J. Theron

Practical mining aspects should be considered when conducting pillar designs for bord-and-pillar layouts. The current methodology for pillar design will result in increasing pillar sizes with depth. This affects the extraction ratio and will result in onerous ventilation requirements when cutting large pillars. A holistic approach, including all mining engineering requirements, is required to ensure that the rock engineering designs are optimized to ensure efficient mining operations and sustainable production. Bord widths should not only be a function of the rock mass ratings, but should also be selected to fit the specifications of the mechanized equipment. The use of a ‘squat pillar’ formula for hard rock is discussed in the paper and the formula based on the exponents of the Hedley and Grant pillar formula, is explored. The effect of abutments and geological losses on average pillar stress is also explored. These factors must be considered when designing layouts at increasing depths.

Nov 27, 2024

By W. K. C. Boemer, G. A. Davies

Vanadium slag production in the shaking ladles started in April 1968, steelmaking was begun in July 1968 and the first continuously cast blocks were produced soon after on the 29th July, 1968. Nominal monthly production of over 20 000 tons was reached in seven months and by December 1970 about 12 000 heats were cast and a monthly production of 35000 tons has been achieved.