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By C. Musingwini, T. C. Sebutsoe

"The primary objective of any mining business unit is to make profit by extracting, processing, and selling minerals from a particular mineral deposit. It is important to optimize extraction of the mineral resource given time, capital, and space constraints. Mineral extraction is often associated with uncertainty due to variable technical and human factors. Technical factors such as grade distribution, ground conditions, and equipment reliability influence the performance of a mining production system (MPS). The performance of the MPS is also impacted by human factors such as employee skills, health, and attendance. Uncertainty associated with technical and human factors often leads to actual output differing from what was planned. Therefore an in-depth analysis of the significant causes of deviations from the planned outcomes is relevant.In this paper we investigate the empirical relationships between inputs and outputs in a MPS in order to assist management in directing efforts at key production drivers. Once a MPS has been empirically characterized, more effort and resources can be focused on the key decision-making variables (DMVs) in order to meet the planned outcomes.For a typical underground, tabular, narrow-reef conventional drill-andblast platinum mine the face advance, face length mined, number of teams, team efficiencies, and team size have a statistically significant relationship with the centares (m²) produced, which is a key performance indicator (KPI). The production function derived in this study can be used to align the physical, technical, and human factors together to predict the optimal output level. The most significant production lever of the MPS is the face advance. IntroductionA mining production system (MPS) is a result of an iterative process of design, planning, and optimization of mining input variables and decision-making variables (DMVs). The MPS exists within the mineral extraction link of the complete mining value chain. It represents the stage where mining companies have the opportunity, through production, to start generating returns on the investments undertaken. Returns on shareholders’ investments can be realized at this stage."

Jan 1, 2017

By Karl-H. Daum

Over many decades, numerous feasibility studies have demonstrated, that the production of sulphuric acid remains the most viable option of sulphur recovery from smelter off gas and abatement of SO2 emissions to the atmosphere. This is particularly more pronounced as smaller but more concentrated off-gas flows are to be treated from smelters, and enhanced sulphuric acid processes be-come available. Off-gas handling systems represent a significant capital and operating cost burden to the metallurgical operation. Modern pyrometallurgical smelter processes for sulphide ores based on the use of oxygen-enriched air, produce relatively small off-gas flows with high SO2 concentrations in the smelter gas of 30?60 %-vol. of SO2. This is a prerequisite for substantial cost reductions in the smelter off-gas handling and treatment system. As an alternative to sulphuric acid production, numerous scrubbing concepts with alkali or dual-alkali combinations as well as organic absorbents have been proposed. Also the reduction of the SO2 to elemental sulphur has frequently been studied. Very few of those alternative processes have been built in industrial scale, but all were generally characterized by none sustainable operation due to cost reasons, problems with issues related to chemicals used and by-products or poor availability. Thus the traditional concept of converting the SO2 to sulphuric acid is most common, although regarded as un-economic, but is at least a proven, environmentally sustainable and reliable way of sulphur gas processing. With the high acid price levels these days, operating companies are even able to generate significant revenue with their otherwise ?fatal? acid. Even though smelter gas is available at high SO2 concentrations, in a conventional acid plant one would add large amounts of air to dilute the gas down to a suitable concentration of 12?13 (14) %-vol. SO2. Dealing with higher SO2 concentrations in smelter acid plants is not feasible, as the gas exit temperature of the catalytic oxidation step would exceed the allowable limit for the catalyst. Hence those conventional acid plants are characterised by the use of large equipment, i.e. high capital cost, high energy consumption and limited flexibility. The newly developed LUREC® process can handle off-gas with significantly higher gas concentrations, even in excess of 25 %-vol. SO2. It is entirely based on well-proven equipment and unit operations. The first industrial application, operating with 16?18 %-vol. SO2 will be presented in this paper, along with the fundamentals of the process. It will be demonstrated that the process requires inherently lower capital cost, fewer operating cost, offers better energy recovery and lower emissions as compared to conventional design. While most existing smelter acid plants would have some built-in spare capacity, any significant increase of smelter capacity, say of 30%, can basically not be accommodated without installing an additional parallel new acid plant unit. An add-on LUREC® module can remove any such restriction, while simultaneously de-bottlenecking the existing acid plant. The paper also discusses the present status of the add-on technology for smelter acid plant expansions with respect to technical boundary conditions of different process alternatives as well as economic and environmental aspects. Application of the LUREC® process to modern smelter acid plants operating with high SO2 gas concentration, will lead to a substantial reduction of the specific size of the gas processing equipment and equivalent savings in capital and operating cost. The LUREC® process is patented worldwide by OUTOTEC.

Jan 1, 2009

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

Jan 1, 1991

Primary roof support represents the first line of defence against rockrelated falls of ground in underground mines, and improper utilization or misunderstanding of the applicability and behaviour of primary support can be costly from a safety standpoint. This is a major concern for underground mines, as roof support is the single most costly expense from a mining operational perspective. This is further backed by the evidence that, in the USA, hundreds of injuries and fatalities still occur each year because of rib, roof, and massive roof falls. Additionally, the fully-grouted passive rebar, fully-grouted tension rebar, and resin-assisted mechanical anchor bolts, which constitute a large portion (89%) of the 68 million bolts installed each year in underground mines in the USA can vary in cost quite dramatically. To mitigate this concern a study was conducted in 2010 by the National Institute of Occupational Safety and Health, in conjunction with Southern Illinois University of Carbondale, to assess the performance of primary roofbolts in underground coal mines for improved safety and cost. This was accomplished using underground roofbolt monitoring solutions, field data, and numerical modelling to better understand the quasi-static behaviour of underground coal mine roofs and the response behaviour of the bolts. In particular, over 170 instrumented extensometers, closure meters, shear meters, fully-grouted passive rebar, fully-grouted tension rebar, and resin-assisted mechanical roofbolts were installed at three coal mines across the USA. Of these three mines, two used the room and pillar extraction method and the other used the longwall extraction method. There was no evidence to indicate a difference in performance of the active primary roofbolts compared with the passive primary roofbolts. Additionally, in the initial loading phase, the active bolts showed no difference in loading, indicating that tension bleed-off is of more of a concern than originally thought. Lastly, for the initial computer modelling studies, challenges still remain in obtaining a good match to the in situ bolt measurements and replicating the discontinuous roof rock and in situ bolt behaviour over time.

Jan 1, 2014

By S. A. Ziebold

There is often a need to increase capacity through existing mist eliminator installations without significant tower modifications. Often there is not enough extra space outside existing vessels unless more expensive changes can be justified. Also, plant downtime must be considered. The challenge in retrofitting is to overcome velocity limitations inherent in the design while at the same time extend performance capability. The solution most often is to create new ways to add extra filter area while remaining constrained to the existing space. This strategy, however, can dangerously backfire if basic design principles are not followed. Mist eliminator design is a complex integration of many conflicting performance parameters. Often the objective is to increase scrubbing capacity while maximizing collection efficiency and service life and minimizing operating pressure drop (energy cost), particle regeneration (re-entrainment) and overall maintenance costs. Results of plant retrofitting using mesh pads, impaction fibre beds and diffusion fibre beds are presented based on actual plant projects. Often the retrofit project can be more effective by increasing the scope and combining other new plant equipment with the upgrade such as new tower components. In some cases, the type of mist eliminators can be upgraded from impaction type (velocity dependent) to Brownian diffusion type (velocity independent) if additional downstream protection or reduced emissions are desirable, especially at turndown rates. Trade-offs between different fibre bed types and styles are discussed as they relate to performance and maintenance.

Jan 1, 2007

By E. Dadzie

This paper will review some methods used to manage geotechnical risk at Damang Pit Cutback. The pit depth is designed up to 250m with push back on the eastern and western walls of the existed and depleted Railway Cutback (RCB) pit. The geotechnical design was presented by SRK Perth, focussing on the optimisation of the bench face angle which finally evolved with ?steep slope strategy?; 24m at 800 and 24m at 750 on the eastern and western walls respectively. The east wall was designed strategically to optimise the proximity of the tailings dam to the final pit crest of the push back. The pit has experienced failure in the weathered profile transgressing through the transitional material into the fresh rock at different location of the east wall since development in July 2005. To ensure stability and to minimise ore loss, the east wall has been redesigned using 700 batter angles and reduced berm width of 4m compared to 800 and 8.5 berm width of 24m bench height. The paper will also discuss how slope failures to the east wall had been sustained, managed and monitored to enhance production. Emphasis is placed on slope management at Damang which includes the principles of wall control blast, visual inspection, slope depressurisation and the state-of-the art of slope monitoring using Autoslope automated prism monitoring. Future slope monitoring is considering the use of 3D Laser and Radar to provide early warning so that evacuation could be effected. The intension is to provide mine personnel the means to identify risk areas during mining activities.

Jan 1, 2008

By W. J. Kempson

As a result of the rising electrical energy costs in South Africa, a method was sought to reduce the overall electrical consumption of typical shaft systems. A typical shaft configuration was analysed and the primary energy consumers were identified. The ventilation fans for this system were found to consume 15% of the total energy of the shaft system. It was calculated that more than 50% of this energy is consumed by the shaft itself, more specifically, by the pressure losses that occur in the shaft as the ventilation air passes through it. In order to ensure that the theory being used for the evaluation of these shaft systems is accurate, a total of five shafts were instrumented and the actual pressure losses over the shafts plotted against time. These shafts were then analysed from a theoretical perspective. Finally, in order to ensure a thorough understanding of the behaviour of the ventilation air in shaft systems, the systems were simulated using computational fluid dynamics (CFD) techniques. There were significant discrepancies between the theoretical analysis and the CFD simulation during the initial comparisons. This discrepancy lessened as the complexity of the CFD models increased, until when the complete shaft was modelled using the full bunton sets, the pipes, and the flanges, the difference between the theoretical evaluation and the CFD simulation was small. This result demonstrates that the theory is insufficient and that the interrelated effect of the buntons and fittings has not been fully appreciated by current theory. The final phase of the work presented here was to evaluate the costeffectiveness of using different bunton shapes and shaft configurations. It is shown that the increase in the pressure losses and therefore the direct operating costs of the shaft can vary by as much as 80%, depending on the bunton configuration chosen. The placement of the piping in the shaft can increase the pressure losses, and therefore the direct operating costs of the shaft, by as much as 12%, depending on the placement of the piping in the shaft; this effect includes the use of flanges. The use of fairings on a large cage can reduce the resistance that the cage offers to the ventilation flow by as much as 30%. This, however, does not translate into a direct saving because as the cage moves through the shaft, the overall effect is transitory. These savings can be significant when the items highlighted in this work are applied correctly. Keywords shaft design, ventilation, energy consumption, modelling, CFD.

Dec 1, 2013

By D. G. Krige

D. G. Krige: The stimulating contributions from Messrs Cooke, Curtis, Martus, Johnston, Van Wyk, Van Rensburg and Wroth are appreciated by the author and provide clear evidence of the importance not only of the principle of assistance to gold mines, but also of its scope and form. In general all the arguments, raised by Messrs Cooke, Martus and Johnston in favour of extending the scope of the new assistance scheme can be accepted as sound.

By I. du Plessis

The metallurgical response of a difficult to treat coal and an easy to treat coal orebody were investigated in terms of the coal yield and the coal quality in a basic coal recovery flowsheet. The purpose of this paper is to show trends and derive principles for the treatment of different coal types. This investigation entailed using the LIMN simulator to simulate the metallurgical response by varying the plant operational parameters. The controllable plant variables investigated were: ?Cyclone size ?Screen and hydro-cyclone cut points ?Medium cut density ?Feed particle size distribution.

Jan 1, 2010

Geological mapping makes an important contribution to orebody knowledge during the evaluation of structurally controlled gold veins. Orebody knowledge comprises an understanding of mineralization characteristics relevant to evaluation, mining, processing and environmental issues and allows project expectations to be realistically set. Structural interpretation to define continuity and mineralization controls is a key part of geological risk reduction and the reporting mineral resources using the JORC and other codes. The Dolgellau gold belt is a small British field that produced an estimated 150 000 oz from narrow quartz veins. Historical records, recent exploration and production data, published material, and accessible surface and underground exposure provide a valuable case study. Previous mapping has shown the general geometry of vein systems, but tends to accentuate gross geological continuity. Recent work has attempted to determine internal vein architecture and examine the significance of the historical views of ore shoot controls. The work resulted in the definition of vein continuity scenarios for the gold belt and evaluation expectations associated with each. This has, in turn, led to a realistic view of what can be achieved from an evaluation and project development perspective.

Jan 1, 2009

By N. C. Gay

The quantitative evaluation of the performance of fill materials placed in gold mine stopes is achieved by (i) measurement of the in-situ load- displacement, or stress-strain, characteristics of the materials; (ii) an objective comparison between the observed fill performance and that predicted from suitable computer modelling techniques; and (iii) objectively comparing conditions in backfilled and conventional stopes. Considerable problems have been experienced with the development of sufficiently robust instrumentation which can be installed within filled pad- docks and give reliable information over a long period of time. However, at most mines where backfill performance is being monitored, it has been possible to correlate stress measurements within the filled paddocks with strain measurements adjacent to the paddocks. The resultant stress-strain curves agree relatively well with those determined in the laboratory. Measurements in the hangingwall strata of filled and unfilled stopes indicate that the placement of fill does reduce the amount of inelastic convergence due to bed separation. Moreover, in some stopes, the rate of convergence with respect to face advance does reduce to that of the theoretical elastic convergence. Quantification of changes in the conditions at the stope face due to placement of fill has proved difficult. Observations indicate that, provided the fill is placed close to the face (- 5 m), and adequate temporary face support is used, conditions in filled panels are superior. In addition, in filled panels the damage caused by a rockburst in the stope and, particularly, in gullies is generally much less than would be expected in an unfilled stope. To quantify these improvements, parameters such as stope height, face advance and the frequency of falls of ground are being monitored.

Jan 1, 1988

By D. Vogt

South African underground hard rock mines are typically managed using measurements made daily, weekly or even monthly of key parameters such as face advance, readiness to blast, blast success, temperature or dust levels. Safety and health can be greatly improved if a real-time measurement system can inform decision making. To enable this vision of widespread sensing, communication and decision support, CSIR has developed an open standard architecture for communication of sensor data, and a reference implementation using that standard. The standard is called AziSA, which means ?to inform? in isiZulu. AziSA provides an architecture that allows for connection of any type of sensor and that is particularly suited to wireless sensing. On the safety side, the CSIR is developing a suite of sensors specifically around the risk of rockfalls. Sensors such as an electronic replacement for the barring tool, or a thermal sensor that can detect loose rock, can already be used to identify potential hazards in the hangingwall. When these sensors are combined with location and time information in a single database, it becomes possible to build maps of risk and to extrapolate risk into unmined areas. It is also easy to confirm that routine safety procedures like barring are actually taking place. In the future, it is proposed that entry inspections will be made using remote techniques that will not place an individual miner at risk. While miners will still be required to bar or place support, they will do so with a clear idea of which portions of the hangingwall are safe, and which present risk. On the health side, if a network such as AziSA is deployed in the stope, it becomes much easier to monitor the condition of individual miners. The work strain experienced by miners can be monitored through heart-rate sensors, and their core body temperature can be monitored using various novel techniques discussed in the paper. Additional sensors can be added to quantify worker exposure to hazards such as noise or dust. Individual sensors allow for people who are at an unacceptably high risk of developing heat disorders to be treated timeously, and they also allow for team management based on objective measures. Over a longer period, position sensing and environment monitors, or personal dosimetry, provide a powerful management tool to prevent workers from being overexposed to hazards, and to confirm that management instructions are being followed. Cost-effective sensing in the stope is a major challenge, but one that can be overcome through technology like that used in AziSA. Sensing provides opportunities both to make immediate interventions when workers are exposed to unacceptable risk, and to manage the long-term exposure of workers to hazard.

Jan 1, 2010

By L. Capponi

At Tapira Mine (Tapira/MG/ Brazil), the largest phosphate concentrate producer in Brazil, the mining is fully mechanized and carried out using open pit mining. This work presents a review of the sampling protocol used for short term mine planning at different stages. Sampling is carried out in two steps. Firstly, diamond drill cores are sampled at 50 x50 x5 metre pattern. Fosfertil analyses grades of P2O5, Fe2O3, CaO, MgO, SiO2 and TiO2 plus geological logging. This information feeds the long term block model. Secondly, drill cuttings are collected at the mining faces released for blasting. The drilling pattern is staggered and when blasting is not necessary, a drill hole is made centered in the block area (25 x 25 x 5 m). The primary samples have mass approximately 60 kg and the same chemical parameters are analysed. Metallurgical and mass recovery tests are performed as well. These samples are used to build the short term block model. At the final stage, the short term mine planning model is used to build a stockpile blending model which is simulated using a spreadsheet to match the goals established by the process plant.

Jan 1, 2009

By K. Andres

This paper presents the vibrating mill technology and summarizes the grinding principle of ultra fine grinding. In addition, a variety of operations is described and the benefits of these different operating modes? product size and efficiency are specified. A case study of an industrial application is presented to show the integration of vibrating mills into grinding circuits. The primary focus is the overall process scheme of this technology, the installation and operating costs, as well as the design parameters required for best performance. The vibrating mill can be used in a wide range of applications for process engineering duties. For a wide range of applications, from soft to extremely hard products, which require ultra fine grinding, surface activation or homogenization, the vibrating mill is well proven. Due to its easy operation, versatility and operation efficiency, the PALLA? vibrating mill gains in importance in the mineral processing industry. Keywords: Coal, comminution, grinding, mineral processing

Jan 1, 2010

By M. F. Handley

This paper covers in situ stresses in the Earth?s crust prior to any man-made disturbances, such as mining. It introduces the Southern African Stress Database, which contains primitive stress measurements obtained from locations spread all over southern Africa. The measured stress data shows large variability without any visible trends, except a relationship with increasing depth. The stress database is reviewed briefly, establishing means to measure the variability or dispersion in the measurements, and showing that the dispersion is not as much a result of experimental error as it is a feature of primitive stress. The paper demonstrates from the beginning that the state of stress in rock is highly variable, but that there are well-defined maximum and minimum limits to all the stress components in rock. Formal error analysis is introduced to check the consistency of the database and to separate out a database of consistent stress measurements for use in a primitive stress model. The aim is to provide a picture of primitive crustal stress based on objective stress measurements together with interpretations of how the primitive stress can be affected by the five main influences; namely, depth, rock mass properties, tectonism, isostacy, and erosion. Four elementary models for primitive stress are introduced and compared with the measured data. It is quite clear that none of the elementary models is sufficient to describe the data. In the absence of a better model, this paper suggests a generic model based on the Hoek-Brown failure criterion and the consistent stress database, since it incorporates the variability in the stress tensor that is likely to be encountered underground all over southern Africa. Rock engineers should take every opportunity to obtain local primitive stress data at every mining operation and civil engineering project, and to adjust the proposed model accordingly Keywords primitive stress, Southern African Stress Database, error analysis, generic primitive stress model.

Jun 1, 2013

By F. Riedel

In May 2009 a test work programme was carried out at the sorting laboratory of CommodasUltrasort in Germany which was aimed at establishing the efficiency of X- ray transmission sorting for the recovery of unliberated diamonds in a size range of +8mm. A PRO Tertiary XRT belt sorter with a working width of 600mm and a belt speed of -3mJs was used for the test work. Dual energy X-ray transmission sorting is a type of sensor-based sorting which uses an X-ray scanning system, The attenuation of the transmitting X-rays is measured at two different energy levels, Thereby it is possible to eliminate the effect of the particle thickness on the measurement result. The effective measurement results in a classification by elemental order based on the periodic table.

Jan 1, 2010

By Johanna van Deventer, Sean Scott, Rene du Preez, Steven R. Izatt

In order to meet the stringent specifications associated with the production of LME Grade A cobalt cathode, producers need to ensure that a number of impurities, including Cd, are removed from the advance electrolyte to ppb levels (~0.1 mg/L). Conventional options include ion exchange using amino methyl phosphonic acid resin and solvent extraction using D2EHPA. More recently a Cd-selective resin, employing Molecular Recognition Technology (MRT), became available from IBC Technologies, Inc. These three potential unit operations were evaluated in parallel at Mintek. MRT was identified as the most technically feasible option. Results obtained on a cobalt advance electrolyte, containing 55 g/L Co, 6 mg/L Cd and 60 g/L Na2SO4, are presented in this paper.

Jan 1, 2007

By V. H. R. Brereton

The gold deposits of the farms Lisbon and Berlyn in the Eastern Transvaal by V. C. H. R. BRERETON* (Visitor) This paper describes the discovery and recovery of gold, over a period dating back to the 1870s. on the farms Lisbon and Berlyn, situated approxim ately 1I km north-east of Pilgrims Rest. where mining ceased in July 1971. 99 years after gold was first discovered in the Pilgrims Rest and Sabie districts. Included in the paper are accounts of the discovery and mining operations of the Lisbon-Berlyn Company from 1884 to 1890. The operations of small companies and syndicates. the superficial winning of gold by diggers, and the encouraging reports of four mining engineers in 1890 are also mentioned.

Jan 11, 1973

By M. Yavuz

Determining the most convenient plant location is one of the most commonly encountered problems in engineering applications. This paper presents a fuzzy multiple attribute decision making (FMADM) model which is developed for the selection of the optimum plant location for natural stone factories in the mining industry. The criteria affecting the decision making process in natural stone industry were determined to solve the plant location problem in the FMADM model. To determine the optimum plant location for a new natural stone factory, which is planned to be established by a mining firm located in the Eskisehir region of Turkey, an analysis was carried out by incorporating the method and the analytic hierarchy process (AHP) method. The analysis shows that the Yager?s method can easily be applied to the selection of plant location in the natural stone and mining industry when the difficulties encountered in the application process of the AHP method is taken into account. Keywords: Natural stone, plant location, decision making, Yager?s method, analytic hierarchy process (AHP).

Jan 1, 2008

Geologists in some underground gold mines collect grab samples from broken ore piles or trucks as a method of grade control. It is often known as muck sampling. Generally, the goal of grab sampling is to try and reconcile the mined grade at the ore source to the predicted grade and/or predict the mill feed grade. The mass of the sample collected is limited by health and safety issues, as well as the capacity of the laboratory to process the samples within a given timeframe. In general terms, grab sampling is known to be problematic because samplers tend to over sample the fines, and/or pick out high-grade fragments; surface sampling of piles does not test material within the pile; muck piles in development drives/faces are likely to be zoned due to the blasting sequence; high or low-grade material may preferentially segregate in the pile during mucking; the five percent mass reject size of the material in muck piles is very large from underground blasting; some correlation usually exists whereby the larger fragments are enriched or depleted in the critical component of value; and the average error made in estimating the true stockpile grade is likely to be high. The method is prone to chronic fundamental sampling, grouping and segregation, delimitation, and extraction errors. Substantial warnings must be given with respect to the use of grab sampling for grade control in gold mines. The method may appear to work sometimes, which can be attributed to a fine gold particle sizing and more disseminated distribution. As with all sampling methods, its appropriateness must be determined by ore characterization and heterogeneity testing to ensure the method suits the ore type.

Jan 1, 2009