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By H. Lieberwirth, M. Klichowicz

The comminution at grain size level for the liberation of valuable minerals usually requires the highest size-specific energy. Therefore, the full understanding of the comminution process at this level is essential. Models based on the Discrete Element Method (DEM) could be a helpful tool for this purpose. In those, the mineral microstructure, which has a dominant influence on the fracture behaviour, can be described in great detail. One major concern, however, is the missing representativeness of simulated models. In this study, a method to overcome this limitation is presented. The authors show how a realistic microstructure can be implemented into a particle bed comminution simulation using a two-dimensional Bonded Particle Model (BPM). The improved algorithm based modelling approach is exemplarily compared to an equivalent real experiment. Thereby, the focus is set on the analysis of the interfacial breakage and the capability of the simulation to reproduce the experimental results. The comparison of the results shows that simulations at grain size level allow a more detailed evaluation of the fracturing behaviour. This is of special interest as the aim of comminution in mineral processing is not only the size reduction of coarse particles, but an efficient liberation of the valuable components. Hence, simulations with automatically generated real mineral microstructures will help to further improve the efficiency of ore processing. Keywords: Discrete Element Method, grain modelling, particle bed comminution, realistic synthetic

Jan 1, 2020

By A. A. Rogozhin, S. I. Anufrieva, A. V. Kurkov

The results of development of technologies for flotation concentration of two types of complex beryllium ores are presented: beryllium-fluorite in apocarbonate metasomatites (Type I) and rare-earthberyllium in feldspar metasomatites (Type II). The researched samples of Type I ores contain a mineral complex that is difficult to separate by fatty acid flotation, including phenakite, bertrandite, fluorite, and carbonate minerals. The ores are characterised by an increased content of fluorite and carbonates. The sample of Type II ore contains such rare-earth minerals as bastnesite and orthite, and a beryllium mineral - cleiophane. The size of individualised grains of cleiophane does not exceed 0.04 mm, which significantly limits the possibility of extracting the mineral into the enrichment concentrates. For the Type I ores, a technological process for the extraction of fluorite using a phosphorus-containing reagent from the class of dialkylphenyl polyethylene glycol esters of phosphoric acid (DAPEGEPhA) has been developed. The use of DAPEGEPhA enables the production of a high-grade fluorite concentrate (95% CaF2). The production of commodity phenakite-bertrandite high-grade (10% BeO) flotation concentrate from the fluorite flotation tailings with using a fatty acid collector is based on binding calcium in the flotation pulp using soda, sodium hydroxide and sodium tripolyphosphate. The use of a combined collector based on fatty acids and DAPEGEPhA made it possible to produce a collective rare-earthberyllium- lithium concentrate with high REE extraction (88%) from the Type II ores. The concentrates obtained are suitable for production of commercial-grade hydrometallurgy products. Keywords: Beryllium ores, flotation, reagents, concentrates, hydrometallurgy products

Jan 1, 2020

By F. Nakhaei, M. Le Roux, Q. P. Campbell, S. Botlhoko

The results of coal sink and float tests are plotted as washability curves. Analysis of these curves indicates the most effective method by which the coal can be cleaned. Sink and float experiments are samplingdependent, destructive, and use toxic chemicals. The present research introduces an alternative method based on a 3D imaging system called RhoVol, in which the density of individual particles is determined from mass and volume measurements. A quantitative error analysis was conducted on coal sample density measurements obtained by RhoVol, and the main factors influencing the errors identified and investigated. The results show that the drawback of this technique is related to its not being able to detect hidden concavities in the particle, resulting in an overestimation of particle volume. Furthermore, it is difficult to capture the impact of porosity and cracks on the volume of coal particles obtained from 3D silhouette images. This mismatch between the RhoVol data and sink and float test results has led to attempts to apply neural network (NN) and linear regression techniques to produce a fully reliable model for correction and estimation of coal density. The NN approach offers superior predictive capability over linear regression, and the estimated density distribution is in line with the sink and float analysis. The root mean square error in estimation of density using the NN model was less than 0.05 g/cm³.

Aug 2, 2022

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

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

Apr 14, 2024

By A. Cambitsis

We show that the prioritization of business improvement and performance management decisions may be significantly improved by taking into account the systemic effects of changes made in a single process area. To do so we build a model of a mining value chain, taking into account consequential downtime that arises from downstream and upstream starvation and choking effects of each activity. Introducing the concept of internal capacity, this approach yields insight into the relative importance of key metrics such as loading and hauling rates. We suggest that a holistic view of this type should be a core component of business improvement and performance management decisions.

Jan 1, 2012

By B. S. Xakalashe, M. I. Khama, Q. G. Reynolds

The carbonaceous reduction of manganese ore is an energy-intensive process that releases large amounts of pollutant gases. In an attempt to circumvent the greenhouse gas emissions, this research investigates hydrogen as an alternative reductant. The reduction of manganese ore with hydrogen results in the structural changes in the intermediate products, which might limit the diffusion of the reactant gases through the pores. Computational fluid dynamics (CFD) models capture the structural changes and predict the rate controlling mechanism and this is key in designing the prereduction reactors. The current work uses CFD models to predict the degree of reduction in a shaft furnace under hydrogen reducing conditions. Manganese ore is described as a collection of scalar and tensor fields such as thermal conductivity of the porous medium, density, and viscous resistance. The results indicate that advection contributes significantly towards the overall transport of hydrogen to the reaction sites. Porosity of the final product was found to be higher than porosity of the raw material.

Mar 4, 2025

By R. V. Ramani, R. Stefanko

"The favorable mining conditions in the United States of America that resulted in the development of the highly productive room and pillar systems are largely disappearing. Increasing mechanization, poorer natural conditions, and greater concern for health, safety and environment gave impetus to the enactment of the Federal Coal Mine Health and Safety Act of 1969. These regulations prescribe more stringent requirements in face ventilation and mine air distribution, necessitating significant changes in existing ventilation systems. Stringent mandatory standards were applied to respirable dust control for the first time, and simultaneous control of both gas and dust has affected the face as well as primary distribution, most notably on haulage entries.The size and complexity of the present-day mining operations, and the need to evaluate various system changes in a limited time, call for scientific methods in mine planning. In this paper, the authors trace the previous applications of computers to mine ventilation planning and discuss the results of their work in applying a ventilation simulator developed at Penn State. The paper includes an analysis of current trends and suggestions for future research and development."

Jan 1, 2014

By D. M. C. Valbom

The introduction of new diamond sorting technologies as well as additional manufacturers I vendors, has increased the equipment selection and combinations available for recovery flowsheet design. This paper describes the methodology used and the benefits realized in the design of recovery plants by ensuring a fundamental understanding of the advantages and limitations of the technology embodied in the equipment, a comprehensive knowledge of the ore body characteristics, and the effective matching of this information in combination with operational commercial requirements.

Jan 1, 2010

One of the main bottlenecks in the carbon-in pulp process for the recovery qf gold is the elution procedure, which typically requires the heading of caustic cyanide eluants to high temperatures for 16 to 24 hours. The present work demonstrates that sodium sulphide solution as an alternative eluant can effect complete elution in about 4 hours at ambient temperatures. Elution efficiencies of around 100 per cent were obtained in 4 hours with a single pass of eluant containing 0,2 M Na2S and 0,4 NaOH about 10 bed-volumes of eluant The initial rate was slow over the first hour of elution, probably because the activated carbon catalysed the oxidation of sulphide to polysulphide. Elution efficiencies of around 100 per cent were also obtained in less than 4 hours during the batch elution of carbon at liquid-to-solid ratio of about 100. Lower liquid to solid ratio resulted in the re-adsorption of gold, probably owing to the oxidation of sulphide to polysulphide, with the resultant formation of gold complexes that were eluted less readily. Improved rates of elution were obtained at higher sulphide concentrations and pH values greater than about 13. Higher temperatures increased the initial elution rate, but lowered the overall extraction efficiency, probably because of the deposition of elemental gold on the carbon. Variation of the ionic strength by the addition of NaCl had no defect on the elution, which confirms that the elution mechanism in the case of sulphide is different from that when cyanide or I1)1droxide is used as the eluant. It is proposed that the elution of gold by sulphide solutions proceeds by means of an initial step that involves the reaction of polysulphide ions with the adsorbed aurocyanide speeds, forming AuCN on the carbon and thiocyanate in solution. This step is followed by the formation of poorly adsorbed complexes with sulphide ions, such as [AUS3-2]-The presence of polysulphides, whether generated in situ by the catalytic oxidation effect of activated carbon or by the addition of elemental sulphur, reduces the elution rate and efficiency dramatically. This is probably due to the formation of complexes such as AUS-3 and AUS-2 , which have a high adsorption affinity.

Jan 1, 1994

By H. M. Stewart, N. F. Dawson, J. P. Beukes, P. G. van Zyl, Q. P. Campbell, A. Jordaan, R. I. Glastonbury, L. N. Sadiki

"SynopsisThe physical properties of oxidative sintered pellets produced from typical South African UG2 ore are compared with the physical properties of pellets produced with conventional South African metallurgical-grade chromite ore (from the Lower Group 6 or the Middle Group 1 and 2 seams). A statistical evaluation of the cured (sintered) compressive strengths proved that pellets prepared from UG2 ore are likely to have the same, or better, compressive strength than pellets prepared from metallurgical-grade chromite ore. The cured abrasion strength of the UG2 pellets was also superior to that of the metallurgical-grade pellets. Scanning electron microscopy (SEM) backscatter, secondary electron, and elemental X-ray mapping were used to determine the reasons for the general superior strength of the UG2 pellets. The case study UG2 ore also required 13 kWh/t less energy for milling to attain the required particle size distribution prior to pelletization, which can lead to substantial cost savings. Results presented in this paper can be utilized by ferrochromium (FeCr) producers to better quantify the advantages and disadvantages associated with the use of UG2 ore for FeCr production.IntroductionChromite, a mineral of the spinel group, is the only commercially viable source of virgin chromium units. Geologically, commercial chromite deposits in the world are found in three forms, i.e. alluvial-, podiform-, and stratiform-type deposits (Murthy et al., 2011; Cramer et al., 2004). Alluvial deposits were formed by weathering of chromite-bearing rock with the subsequent release of chromite and gravity concentration by flowing water. However, these deposits are relatively small and of comparatively minor commercial interest. Podiform-type chromite deposits usually occur as irregularly shaped pods or lenses, and their distribution within a mineralized zone is usually relatively erratic and unpredictable, making the exploration of these deposits difficult and costly (Cramer et al., 2004). Stratiform-type chromite deposits occur as parallel seams in large, layered igneous rock complexes with more regular layering and lateral continuity (Cramer et al., 2004). The largest example of a stratiformtype chromite deposit is the Bushveld Complex (BC) in South Africa, which holds an estimated three-quarters of the world’s viable chromite ore resources (Cramer et al., 2004).Several chromite seams exist in the BC. The seams of economic interest are the Lower Group 6 (LG6), the Middle Group 1 and 2 (MG1 and 2), and the Upper Group 2 (UG2) seams (Cramer et al., 2004). The LG and MG seams are specifically exploited for their chromium content, while the UG2 seam is mined primarily as a source of platinum group minerals (PGMs) (Xiao and Laplante 2004). Extraction of the PGMs from the UG2 ore usually involves the liberation of the sulphide by milling and subsequent recovery of the PGM concentrate through flotation (Xiao and Laplante, 2004). Chromite in the PGM concentrate is undesirable; therefore, PGM recovery circuits are specifically designed to ensure maximum rejection of the chromite to the tailings stream. However, the chromite in the PGM tailings can potentially serve as feedstock for ferrochrome (FeCr) production, after beneficiation to increase the chromium content."

Jan 1, 2015

By E. Moosavi, K. Tolouei

In mine production scheduling problems are well-known, where most are categorised into deterministic and non-deterministic NP-hard problems because of their complexity. Many researchers have attempted to solve the problem by applying various optimisation techniques. While using traditional methods they observed extreme difficulty in solving highly complex problems. In the 1990s, many researchers addressed long-term production scheduling (LTPS) by using simulated annealing, the genetic algorithm, the tabu search algorithm, and the ant colony algorithm. These algorithms are known as the meta-heuristic algorithms, and have proved to be the most efficient algorithms for solving LTPS thus far. The firefly algorithm is one of the best methods for visualising problems related to LTPS, and provides the best possible optimisation. It is one of the simplest methods, and easy to apply to any NP-hard problem. The experimental, and comparison with other recently published algorithms shows that the proposed algorithm is feasible, and is an effective approach for long-term production scheduling problems. The results indicated that a better solution is obtained by the proposed method, compared to other methods in terms of cumulative net present value, and average ore grade. Additionally, the CPU time via the proposed model is roughly 2.52% higher than the other method.

Sep 1, 2021

By A. H. Mokken, G. K. I. Blendulf, G. J. C. Young

A study of the arrangements for pulp discharge on pebble mills, and their influence on mill performance by A. H. MOKKEN, B.Sc. (Eng.) (Rand) (Fellow), G. K. I. BLENDULF, M.Sc. (Stockholm) (Member), and G. J. C. YOUNG, B.Sc. (London) (Member) Although used for many years in the gold-mining industry, the conventional pan lifter at the discharge end of a pebble mill has hindered attempts at optimization of speed of rotation and pulp flow. A theoretical approach to flow patterns in pan lifters has pointed to limitations in the capacity of these lifters, with capacity decreasing as speed of rotation is increased. Complete elimination of the pan lifter by the adoption of peripheral or end discharge has resulted in a marked increase in pulp flow and an improvement in power utilization. Assays of the gold content of pulp discharged after a mill stoppage from mills fitted with pan lifters show values well in excess of the current milled grade, which indicates an accumulation of gold particles in the mill and, therefore, a longer retention period. The same could apply to other heavy minerals and metals present in the ore feed or worn off mill linings and metallic grinding media. Examples are osmiridium, fine iron and iron alloys, and copper and sulphide minerals. This phenomenon is attributed to the higher specific gravities of these metals and minerals, causing them to migrate towards the mill shell, from which position they are not readily removed by the lifters at the discharge end of the mill. The close association of gold particles with iron and sulphide minerals in the crevices of mill liners might be conducive to the formation of coatings on the gold. Assays of samples of milled pulp taken from a mill designed for end discharge show a lower accumulation of gold. The accumulation and nature of these gold particles are being studied at present. With increasing diameters of pebble (or autogenous) mills, peripheral or end discharge appears to be the preferred method of pulp removal.

Jan 5, 1975

Numbering of the Journal Because computer and library authorities prefer journal subscriptions to be based on the calendar year, and because the present numbering system of the Journal of the South African Institute of Mining and Metallurgy does not conform with the calendar year, it has been decided to change the numbering system of the Journal.

Jan 1, 1979

By T. Bednarski, L. Thomas, G. Banza, D. Muteba, P. Cole, M. Soderstrom

"In copper solvent extraction, aqueous-in-organic entrainment results in the carry-over of unwanted species into the electrowinning process with negative cost implications. The need to bleed is increased and the purity of the copper cathode is potentially compromised, depending on the impurities being transferred (Cl, NO3, Fe, Mn). In some cases, entrainment results in elevated levels of manganese in the electrolyte which can cause oxidative conditions to develop and degradation of the organic phase. Degraded organic phase detrimentally affects the physical performance of the copper solvent extraction process, contributing to even higher aqueous-in-organic entrainment. Aqueous entrainment may be minimized through good operating practices and maintenance of the organic quality, but the accurate measurement of entrainment remains a challenge. This paper presents coalescing devices designed to provide more exact measurements of aqueous-in-organic entrainment. IntroductionMany copper solvent extraction (SX) plants experience high aqueous-in-organic (A-in-O) entrainments but are unaware of the extent of the problem because entrainment is often not measured. Aqueous entrainment results in the carry-over of unwanted species into the electrowinning (EW) process and the need to bleed to maintain acceptable concentrations in the electrolyte. The costs associated with the bleed are exacerbated in copper plants where aqueous entrainment levels are high. Other than these direct costs, there are operational consequences of the build-up of certain species in the EW circuit that can have deleterious effects on plant performance.Minimizing aqueous entrainment in operating plants requires an understanding of the factors that affect entrainment levels. Both chemical and operational influences are discussed in this paper. The difficulty of obtaining a representative sample makes it challenging to obtain an accurate measure of aqueous entrainment from a continuous process like SX. Coalescing devices used to get a more precise measurement of A-in-O entrainment are useful to complete mass balances around the SX circuits. Alternatively, mass balances can be used to estimate aqueous entrainments and are helpful as a tool to troubleshoot operations with a view to improving plant performance."

Jan 1, 2016

By J. Petersen, K. Shaik

The traditional route utilised for the recovery of Platinum Group Minerals (PGMs) in South Africa involves the concentration of ore by flotation, followed by smelting to produce a PGM rich matte, which is then digested hydrometallurgically to solubilise base metals and Platinum Group Elements (PGE). HCl/Cl2 is conventionally used for PGE dissolution from concentrated material, which is an aggressive reagent associated with high maintenance costs as well as environmental concerns. Of interest is the leaching of PGMs directly from mineral ores, low grade concentrates and tailings in a cyanide-ferricyanide system as a more sustainable alternative. This system has shown favourable Pt recoveries among the Pt minerals, but with sperrylite (PtAs2) displaying highly refractory behaviour. The dissolution of sperrylite is investigated in a cyanide-ferricyanide system in the context of an electrochemical study involving OCP, chrono-amperometry and cyclic voltammetry. This approach allows the assessment of leaching characteristics as well as the investigation of surface passivation, which is believed to contribute to poor Pt recoveries. Some evidence is presented that such passivation is related to the formation of arsenide species on the mineral surface. Keywords: Sperrylite, PGMs, cyanide, ferricyanide, electrochemistry

Jan 1, 2020

By G. Nino de Guzman, L. C. Woollacott

The isotherms that are usually taken to represent the force 'driving' the adsorption in carbon-in-pulp processes may change from contactor to contactor. This paper reports on an investigation into the extent to which this Isotherm shift is Influenced by the changes in oxygen and cyanide concentration that occur in operating plants. The impact of the isotherm shift on the reliability of simulation results is examined.

Jan 1, 1993

By E. Rex, T. J. Goff, W. S. B. van Beek, P. E. Nel

"The PGM smelting industry is known for operating at significant high superheats. Both the slag and matte phases that separate via gravity need to be tapped out of the furnace in a safe and efficient manner, which can be achieved only through sound tap-hole management practices. This paper gives an overview of the tap-hole management approach followed in the operations at Lonmin’s Western Platinum Smelter.Five circular AC furnaces are operated, with each furnace comprising essentially a crucible and auxiliary equipment. Within the crucible, slag and matte tap-holes are situated at different elevations. The slag phase is tapped at a higher elevation than the matte phase. At the three older furnaces, tapping is done by the conventional method of lancing open the tap-hole and closing it with a dolly. However, at the two newer furnaces, tap-holes are opened and closed with the aid of a drill and mudgun.Tap-hole management can be divided in a number of aspects. In order to achieve a safe opening and closure every time, Lonmin distinguishes between three important aspects - the design of the tap-hole system, the operating practices followed, as well as care and maintenance. This paper briefly describes the tapping channel and cooling requirements for handling superheated matte, together with the difference in design of the matte and slag tap-holes. Mention is also made of the importance of auxiliary equipment, which includes the launders, fume extraction, granulation water, and mudgun and drill units. Improvements made to the matte tapblock design are discussed.The monitoring, alarm, and safety systems related to matte and slag tapping are described. Tap-hole life and different wear mechanisms are mentioned. Tapping can be done safely only if the tap-holes and equipment are maintained properly. Mention is made of the tap-hole refractory repair practices, and the improvements made in order to reduce repair time as well as to extend the time between deep repairs. The importance of the correct tap-hole clay selection in order to nurse the tapping channel after each tap is also discussed.The paper finally touches on some research and development work that was done in order to improve the monitoring of the condition of the tapblock by means of fibre optic temperature sensors"

Jan 1, 2014

By E. J. Walls, W. C. Joughin

"In selecting a suitable method to sink a vertical shaft for underground access, a number of constraints influence the ultimate decision of where and how to develop the shaft, not least among these being safety, development and construction time, and cost. Two additional considerations stand out: these being geotechnical conditions and technology, the latter taking into account existing underground access. Assuming a project for which an existing underground excavation is available, it is tempting to build a shaft sinking project from the outset based on the raiseboring method, which has the potential to be the safest, fastest, and least expensive method provided that geotechnical conditions permit.And therein lies the rub: regardless of project time or cost constraints, when it comes to raiseboring a long (say, greater than 500 m), large diameter (greater than 4.5 m) shaft, the rock mass conditions ultimately dictate what method of shaft sinking will be feasible. Over the course of several studies for a particular project, several geotechnical analyses were carried out specifically for the purpose of developing a shaft by raiseboring. Risk analysis and experience showed that where the rock mass conditions indicated an unacceptably high risk potential, an alternative method needed to be considered, even if this meant increasing both the time and financial requirements.In this paper we present an overview of geotechnical investigation practices for shaft sinking. Decisionmaking thresholds for raiseboring or other methods of shaft sinking are discussed, including probabilities of failure, empirical rock mass classification, basic wedge failure, and back-analysis of a failed case. The design of appropriate support, and analysis of relative safety benefits for various shaft sinking methods, falls outside the scope of this work, and will be presented in a separate paper.IntroductionThe potential for rock mass failure during shaft development and construction affects whether a project is completed safely, on time and on budget. During the course of several studies to establish ventilation shafts for an underground expansion project, geotechnical analyses of rock mass stability and failure potential were carried out. The investigations were guided by the method of McCracken and Stacey (1989) together with insights from Peck and associates (Peck, 2000; Peck and Lee, 2007, 2008; Peck, Coombes, and Lee, 2011)."

Apr 1, 2019

By A. M. Starfield

Discussion J. de V. Lambrechts (Fellow): The author's paper is a brilliant follow-up of an earlier paper by Starfield and Dickson.1 I have no quarrel with Dr Starfield's computerization of a complex problem, but I do not believe that he is getting quite the right answers from his programme. My impression is that his predictions about wet bulb temperature increases in very deep mines are over optimistic; in other words, that it will be hotter than Dr Starfield predicts. This is putting my views in a nutshell. This is not the occasion on which to indulge in lengthy argument about the original paper by Starfield and Dickson, but the present paper is, after all, based directly on that earlier paper and if the one fails, the other cannot succeed. I did level certain criticisms at the first paper and cannot say that the authors' replies were very convincing. I do not think it is a sin to admit that I belong to the old school which believes in thorough field experimentation and practical trials and no amount of mathematical manipulation or physical theorizing, no matter how excellent, can make up for inadequate practical confirmation. This, as I see it, is still the crux of the matter. The original paper by Starfield and Dickson still rests on somewhat scanty practical evidence and I would be much happier if Dr Starfield's computer programme, based on the Starfield-Dickson model, had been checked against a large mass of observations in the practical mining situation. This is what both Wiles2 and myself3 had tried to do previously. What we lacked in mathematics and/or computer aids, I think the present paper by Dr Starfield lacks in practical substantiation. This is no condemnation of the author's paper which, taken by itself, is excellent but I think the final stage is still lacking, namely the bringing together of theory and empiricism in a manner acceptable to all. This may be wishful thinking on my part but I hope, within the next year or so, to come up with a modified Starfield-Dickson model in such a way that the computer answers will agree in the majority of cases with the few hundred field observations which are already on record. It might be a case of applying the proverbial 'Cook's Law' to the Starfield-Dickson model! REFERENCES 1. STARFIELD, A. M., and DICKSON,A. J. 'A study of heat transfer and moisture pick-up in mine airways.' J. S. Air. Inst. Min. Metall., 68, (5), 1967. 2. WILES, G. G. 'Wet bulb temperature gradients in horizontal airways.' J. S. Air. Inst. Min. Metall., 59, (7), 1959, p. 339. 3. LAMBRECHTS, J. DE V. 'Prediction of wet bulb temperature gradients in mine airways.' J. S. Air. Inst. Min. Metall., 67, (11), 1967, p. 595. R. Hemp (Visitor): Dr Starfield's paper has very effectively rounded off one particular aspect of the general problem of heat flow in mines. The ease with which this computer programme can be used to calculate temperature increases in horizontal airways must lead to its wider use in ventilation planning and, in developing this rapid method, Dr Starfield has indeed rendered a valuable service to the mining industry. One could consider further instances of heat flow in airways in which the availability of a rapid computer method would be desirable, e.g. the flow of air down a shaft, where there is an increase in temperature due to adiabatic compression, as well as an increase in virgin rock temperature as the depth increases. However, this particular case would not present any new problems and would merely require an extension of the exisitng work. I should like in this contribution to talk about an aspect of environmental control in mines which, I think, will become more important in the future. It is well known that wet bulb temperatures are subject to fluctuations underground. In some instances, particularly in stopes, the fluctuations, both with time and position, can be considerable. The theoretical work which has been carried out on temperature increases has been aimed at the prediction of mean temperatures, and no account has been taken of fluctuations around this mean. It is questionable whether this approach will, on its own, be sufficient, particularly when temperature increases in stopes are considered. The fluctuations in air temperatures underground arise from two causes. The first of these would be the fluctuations in surface conditions, and here one could list random, diurnal and seasonal fluctuations. The second cause is the multitude of things which vary in a mine and here one could list variations in air flow quantity, sources of evaporation, heat transfer from pump and compressed air columns and, particularly in the stope, variations in air flow patterns. Fluctuations arising from surface temperature variations should be amenable to calculation, and here one envisages figures relating the decay of temperature variation with distance to factors such as air flow quantity. Fluctuations resulting from changes in the mine are perhaps more difficult to tackle theoretically and the best approach could well be to analyse underground observations. In this connection, there is a good case to be made for the increased use of statistical methods in the analysis of underground temperature measurements, and it might be of value to look at current air-conditioning practice. When carrying out cooling load calculations for a particular location it is customary to use design wet bulb

Jan 11, 1969

By A. Navarra, A. Quelopana

The mine production scheduling problem (MPSP) has been studied since the 1960s, and remains an active area of computational research. In extending the concepts of the MPSP, the automated mine may now be regarded as a hierarchical intelligent agent in which the bottom layer consists of distributed robotic equipment, while strategic functionality occupies the higher layers. Here we present a disambiguation of artificial intelligence, machine learning, computational optimization, and automation within the mining context. Specifically, the Q-learning algorithm has been adapted to generate the initial solutions for a high-performing strategic mine planning algorithm, originally developed by Lamghari, Dimitrakopoulos and Ferland, based on the variable neighbourhood descent (VND) metaheuristic. The hierarchical intelligent agent is presented as an integrative conceptual platform, defining the interaction between our new Q-learning adaptation and Lamghari’s VND, and potentially other hierarchically controlled components of an artificially intelligent mine, having various degrees of automation. Sample computations involving Q-learning and VND are presented.

Dec 1, 2021