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By Y. -L. Chiang

Paper written on project work carried out in partial fulfilment of BSc (Eng) (Chemical) The generation and control of manganese sulphide particle size distribution using various mixing configurations were investigated. Specifically, this paper discusses how varying mixing intensities on a macro and micro scale affect the resultant manganese sulphide particle size distribution (PSD) at a constant concentration. The mixing variations were achieved using various Y- and T-mixers as premixing devices into a jacketed, agitated vessel. The reagents used to generate the seeds were aqueous manganese sulphate and sodium sulphide solutions. The results showed that the absence of micromixing resulted in very poor control of the particle size distribution, the particle size and the number of particles produced. This was due to the fact that the macromixing time is much greater than the corresponding precipitation time, resulting in different mixing regions existing within the reactor. From the micromixing experiments, the T?mixer was found to be the most effective mixing device at lower concentrations due to the T-mixer providing a faster mixing time than the other mixing configurations, thereby effectively localizing the supersaturation and hence controlling the precipitation. The results obtained from the direct addition without the extension pipe (i.e. short mixing time) indicated a general increase in mean particle size and decrease in total particle number at reagent concentration of 0.007 mol.dm-3as compared to the corresponding micromixing experiments. The phenomenon may be due to incomplete crystallization at the immediate exit of the mixing device. Further experiments need to be conducted before any conclusions can be drawn about the precipitation mechanisms.

Jan 1, 2007

By E. Saari, B. Musuku, K. Jansson

The quality of the concentrator plant process water is crucial for high performing mineral flotation circuits. Significant metal recovery and concentrate grade losses have been noticed particularly in more closed water loops. The recovery and quality losses have started to emerge mainly with decreased freshwater usage and because of changes in tailings management towards tailings thickening, paste, or dry stacking. These changes result in shorter, closed water loops with a rapid build-up of harmful components with negative impact on the beneficiation process. Consequently, water quality has a direct link to decreased product value and sales returns. Complete cleaning or water blending might not be the right way to achieve high recovery and high-quality products. Instead, the key is to understand what kind of water qualities are needed for the flotation of specific minerals. In this study harmful substances in various plant water streams were removed and/or modified by pilot-scale dissolved air flotation. Batch-scale mineral flotation tests with the untreated and treated water were then performed by using fresh ore samples and the plant’s own laboratory flotation procedure. The results indicate that because of dissolved air flotation, the mineral flotation recoveries were enhanced by 1‒3.3%. Keywords: Mineral recovery, metal recovery, mineral quality, process water, concentrator, beneficiation, mineral processing, flotation, process water treatment

Jan 1, 2020

By J. Hadjigeorgiou, G. Knox

There is a considerable wealth of information on the behaviour of rockbolts under axial loading conditions both in situ and under controlled laboratory conditions. However, there is comparably less data on the behaviour of rockbolts in shear. Several authors have contributed to the understanding of the behaviour of conventional rockbolts’ performance under shear loading. Much of this data have been gathered through experimental programmes conducted on sections of the rockbolt. Considering the continuous coupling between the rockbolt and borehole interface, the approach is justified. Due to the design premise of energy-absorbing rockbolts with discrete couplings, full scale testing is required to quantify the behaviour of the rockbolts. This paper reviews published results obtained from two testing facilities, which are the SINTEF rockbolt pull test rig in Trondheim, Norway and the Epiroc Combinations Shear and Tensile test rig in Johannesburg, South Africa, both capable of conducting full scale testing of rockbolts. Based on this high-quality database it was possible to identify three distinct responses of rockbolts in shear. The behaviour under load is controlled by the presence or absence of an encapsulation medium (resin/cement grout), the reinforcement mechanism, and whether the rockbolt is continuously or discretely captured. As a reference the results are compared to the performance in shear of conventional grouted rockbolts.

Nov 24, 2025

By H. E. Parkinson, Y. P. Chugh

About 90 percent of the coal mined underground in the U.S.A. today is mined by the room-and-pillar system. The productivity of the system dropped from about 14,1t per man-day in 1969 to about 7,3 t per man-day in 1978. Since 1974, the U.S. Department of Energy (formerly the U.S. Bureau of Mines) has been engaged in research to improve the production and productivity of the room-and-pillar system. The effort involves the development of integrated, automated, and remote-controlled equipment and systems including remote-controlled continuous miners, automated miner-bolters, automated extraction systems, continuous haulage systems for use behind miners, and remote-controlled equipment for miners, bolters, and shuttle cars. These systems are discussed briefly, the discussion being related primarily to coal winning, although the concepts could be utilized in room-and-pillar systems in mines other than coal mines.

Jan 1, 1981

By A. L. Hinde, A. A. Bradley, K. H. Stanton, P. J. D. Lloyd, G. K. Schymura

As part of its programme to develop methods for concentrating minerals underground, the Chamber of Mines of South Africa in conjunction with Lurgi evolved and tested a I MW centrifugal mill. The machine is capable, in the autogenous mode, of grinding more than 25 t/h of minus 75 plus 44 mm hard quartzite to a minus 3 mm product containing over 50 per cent minus 75 µm material, and as a ball mill of grinding more than 60 t/h of minus 19 mm quartzite to a minus 3 mm product containing over 35 per cent minus 75 µm material. Extrapolation of computer models of the process suggests that these limits can be comfortably exceeded. Long-term continuous testing is in hand, aimed at proving that this I m in diameter by I m long mill is in every respect the equivalent of a conventional 4 m by 6 m ball mill.

Jan 1, 1982

Council Members 1988/1989, left to right: Row 1 (seated): H.E. James, P.W.J. van Rensburg, Dr H. Wagner, B.C. Alberts, C.E. Flvaz (President), H.G. Mosenthal, R.D. Beck, Professor A.N. Brown, Dr P.R. Jochens Row 2 (standing): P.C. van Aswegen, J.P. Hoftman, D. Wllson, P.M.T. White, K.A. van Gessel, Dr LA. Cramer, J.S. Freer, V.J. Moore, G.A. Brown, R.A. Snodgrass, J.D. Austin Row 3 (standing): Dr G.A. Fourle, Dr J. Lurle, A.M.P. Henderson, R.P. Mohring, J.A. Cruise, J.L. Nel, H. Scott Russell, J.P. Deetlefs, D.A.J. Ross-Watt Inserts, top row: K.R. Beard, N.J. De~ine, R.J. Dlppenaar, G.T.G. Emere, Dr J-P. Franzidls, Professor R.P. King, J.G. McCallum. Bottom row: G.Y. Nisblt, Professor R.P. Plewman, Dr R.E. Roblnson, Dr O.K.H. Steften, P.A. van Wlelllgh

Jan 1, 1989

1. Book review Steels- microstructure and properties, by R. W. K. Honeycombe. London, Edward Arnold, 1981. 244 pp. Rl9,90. (Reviewer: P. T. Wedepohl) 2. New books Decision making in continuous steel casting, by Mark A. Vonderembse. New York, Bowker Publishing Co., 1981. U8 pp. Handbook on continuous casting, by E. Herrmann' Diisseldorf, Aluminium-Verlag GmbH, 1980. 742 pp. Cloth DM 1250, leather DM 1500. Pore structure analysis from a nitrogen adsorption study of porous materials using a volumetric method. User's guide to the FORTRAN programs: PORESl and PORES2, by E. G. K. Woods. Pretoria, CERGCSIR (P.O. Box 395, Pretoria 0001), May 1981. Oil and gas from coal, by Dick Olliver. London, Financial Times (10 Cannon Street, London EC4P 4BY), 1980. £120. The seventies: a review, the eighties, a preview, by W. Hibbard (editor). Oxford, Pergamon, 1981. 120 pp. £12.50. Computers in materials technology, by T. Ericsson editor). Oxford, Pergamon, 1981. 230 pp. £14.50.

Jan 1, 1981

By K. L. Morton, R. .. J. Connelly, P. C. Bouw

Uncertainties in the forecasting of the volume of groundwater likely to enter underground workings present difficulties for mine management in planning and costing the water-related activities of mining. This paper describes a technique that was developed to assess future inflows into a gold mine in the Orange Free State. The approach employed was based on the interactive operation of two separate computer models: a regional-flow model and an inflow model. As a first stage, a regional finite-element model of groundwater flow was set up and calibrated. Predictive runs were made to establish the influence of mining on the water levels from base-line conditions. This step was necessary since observations of the regional water levels were sparse. The predicted water levels were transferred to the inflow model, and the future flowrates were calculated. Both models were calibrated by trial and error until a satisfactory match was obtained between historical records and the values predicted by modeling. In this way, an improved monitoring system to record regional water levels and pressure heads within a mine was established. By the use of information, the initial model predictions can be refined as mining progresses. The models used are based on a non-linear relationship between the area mined and the inflows. A phased behaviour in the rates of inflow increase was noted. In conclusion, the modeling approach can provide mine management with guidelines for dewatering requirements. The confidence levels depend upon the records that have been collected during the mine's history.

Jan 1, 1988

By G. Bueno, L. M. Tavares, T. M. Campos

Iron ore pellet feed pressing has been one of the first and most successful applications of High- Pressure Grinding Rolls (HPGR) in the minerals industry. Vale S.A. has been one of the pioneers in this application, currently with six HPGRs in operation in the pelletising plants of the Complexo de Tubarão (Brazil). In spite of the important advances in the technology, risks still exist when scaling-up the process from small-scale information. The present work takes advantage of the recent advances in the mathematical modelling of the HPGR performance, using the population balance model to evaluate the modified Torres and Casali model prediction in iron ore pellet feed pressing for different scales of operation. Results are presented that compare the HPGR performance in different operational scales for iron ore pellet feed pressing, demonstrating the confidence of this model-based approach for scale-up to full-scale HPGRs used in pressing iron ore concentrates. Keywords: Scale-up, HPGR, pellet feed, modelling, simulation

Jan 1, 2020

By Nelson Sánchez

The Los Bronces Development Project (LBDP) was approved in October 2007 by the Anglo American plc Board at a capital cost of some US$1.74B, and is scheduled to be completed by the end of 2010. The project represents a big step forward for the Los Bronces mine since it will increase the treatment rate by 140%, the mining rate by 64% and almost double copper production during the first four years of the project. This economy of scale will also position Los Bronces as a much bigger and efficient mining operation, better prepared to face adverse market scenarios and allowing Anglo to consolidate its position in base metals. The project addressed numerous scope alternatives such as plant capacity vs mine capacity, cut-off grade strategy to maximise value based on the existing flotation and leaching process, cutback strategy, etc. Main infrastructure aspects of the project are as follows: ?Dump area with 2400Mt capacity for waste and low grade material ?Mining equipment: 1 x electric shovel, 12 x 300t trucks, 1 x drill ?Primary crusher: 60? x 89? ?Three new power lines (144MW capacity) ?Grinding plant at 87kt/day nominal capacity (includes a 4.3km tunnel with a regenerative conveyor belt) ?Grinding plant: 1 SAG mill 40? x 25?, 2 ball mills ?Slurry pipeline (52 km from grinding to flotation plant) with five choke stations ?Flotation plant (rougher, cleaning, regrinding circuits and molybdenum flotation). The history of Los Bronces started in 1867 when it was discovered. In 1978 ENAMI (Government company) sold Los Bronces to Exxon for US$112M. The same year a snow avalanche destroyed San Francisco Flotation Plant. In 1979 the plant was rebuilt at 4.9kt/day and from 1979 to 2002 several plant expansions were made up to 54kt/day capacity. In 2002 Anglo acquired Los Bronces for US$1.3B. The 2002 resource and reserve statement showed reserves at 459Mt at 1.02%CuT and resources at 780Mt at 0.80%CuT. The 2007 statement shows reserves at 1,480Mt at 0.67%CuT and resources at 1,644Mt at 0.45%CuT. Defining the size of a plant expansion is the key challenge that any expansionary project faces. The scope of this paper is a general review of the project, describing all the mining techniques that were used to define the best options that maximise value. In this context, a range of plant capacities (from 58 to 230kt/day) combined with a range of mining capacities (from 83 to 166Mt/a) were evaluated in detail, defining for each option a mine plan, Capex and Opex, and cut-off grade strategy for flotation and leaching processes and the associated Net Present Value (NPV) that was used as the selection criteria. The Los Bronces mineral resources potential provides good growth opportunities for the mine beyond LBDP implementation, offering an exciting invitation to continue studying new initiatives that maximise the value of this deposit.

Jan 1, 2008

By A. R. Javadi

Pyrite (FeS2), chalcopyrite (CuFeS2), sphalerite (ZnS) and galena (PbS) generate hydrogen peroxide (H2O2) when placed in water, with pyrite generating the most amount of H2O2. The order of these minerals for generating H2O2 in water is: pyrite > chalcopyrite > sphalerite > galena. In previous research, the methods of generation of H2O2 in the grinding mill were studied. In this study, data from experiments were included by using MODDE 9 software for six parameters composed of water, grinding media, particle size; adding collector to the grinding, and adding depressant and pH. Finally the optimal condition was achieved. According to the results, most of the recovery is related to the particle 75 μ size and when using processed water, adding 100% collector and without depressant and pH 10. The most recovery of lead in the case of the use of fresh water is in the case of 100% collector, without depressant, pH 10 and 150 μ particle size. Maximum recovery in fraction size 45 μ in the case of water process is 100% collector, without depressant and pH 10. Maximum depressant of pyrite will be the fraction size 75 μ, distilled water, without collector, 100 depressant. Keywords: Sulfide mineral, hydrogen peroxide, flotation, oxidation

Jan 1, 2020

By J. Sehata, J. Tshilongo, I. Govender, S. Pikinini, T. L. Moodley, M. Raedani

In the face of dwindling Merensky ore reserves in the early 1990s, Mintek developed the MF2 flotation circuit to economically recover platinum group metals from the shallow upper group-2 chromitite reef. By aligning the primary and secondary grind to the platinum group metal liberation kinetics, excessive fine chromite generation was avoided, and thus one could consistently produce high-grade concentrates (>200 g/t platinum group metals) with low chrome content (<3% Cr2O₃), well within smelter tolerances. This, coupled with upper group- 2's low-cost open-pit mining, made it particularly attractive to investors and mining houses alike. As a result, deeper lying reefs like middle group and lower group with lower platinum group metals grades and complex mineralogy, were primarily mined only for chromite using density-based separation processes (like spirals), of which the waste tailings are best classified as secondary platinum group metals resources. Today, with global platinum group metal demand rising and high grade upper ground-2 reserves rapidly diminishing, the industry must revisit these middle group and lower group tailings. Mineralogical analysis reveals several challenges: Middle group tailings feature platinum group metals locked within silicate matrices, while lower group tailings contain fine chromitite fragments, and are hosted within non-floating laurite grains, each presenting liberation and gangue issues. Smelter constraints still demand concentrates below 3% Cr₂O₃, requiring solutions that consider both platinum group metals recovery and chrome entrainment. To develop a solution, two lower group chrome spiral tailings samples were obtained from a large chrome mine: Sample A (+425 μm – 106 μm) and Sample B (−106 μm). Initial assays showed Sample B contained higher platinum group metal grades but also higher Cr₂O₃. As such, different blends of flotation feeds were evaluated using ratios of 90:10 (C1) and 80:20 (C2) of Sample A to Sample B, targeting the coarse fraction's lower chrome content while recovering valuable platinum group metals from the fines. This work proposes a practical route for mining houses to unlock additional platinum group metal value from lower group ore spiral tailings. By optimising blending ratios and reagent suites, it is possible to meet smelter chrome and platinum group metal grade thresholds, whilst still obtaining economically viable PGM recoveries.

Mar 6, 2026

By X. Zhang, Z. T. Yuan, C. Zhang, H. M. Zheng, X. M. Wang, L. X. Li, N. Li

In metal mines, comminution is commonly used to liberate minerals for subsequent mineral concentration processes, such as flotation, magnetic separation and gravity separation. However, ores are aggregates of valuable minerals and their disseminated gangue minerals, and these minerals exhibit diverse comminution behaviours. In carbonate-bearing iron ores, hematite, siderite, and quartz are the predominant minerals, of which hematite is the objective recoverable mineral. In this regard, the present study aimed to identify the imperfections in siderite minerals by X-ray computed tomography (XCT), and thus quantitatively characterise their size and shape in terms of Feret diameter and aspect ratio, as well as other parameters. XCT images indicated the protogenetic ubiquity of pores <5 μm in diameter and cracks of 7-8 μm inside siderite particles. In the ground product of artificially mixed siderite and hematite, siderite particles were smaller than hematite particles. Siderite particles had a larger surface area and greater surface roughness than hematite particles of the same volume, which was attributable to the intrinsic holes and cracks in siderite particles. This study provides an insight into the relationship between the inherent imperfections of minerals and their comminution properties. Keywords: Siderite; X-ray computed tomography; particle imperfections; grinding behaviour; sliming

Jan 1, 2020

By S. W. Jacobsz, Y. Narainsamy

Several high-impact tailings dam failures around the world in recent years have placed a renewed focus on the stability of tailings dams and pointed to potential shortcomings in traditional drained design and safety evaluation procedures. A need to consider undrained shear strength in the design of tailings facilities has become apparent. However, there are specific requirements that need to be met before undrained shearing occurs. In South Africa, the last major failure was likely the Merriespruit disaster in 1994, which leads to the question of whether the conditions required for undrained shearing are readily applicable to South African tailings dams. This paper describes research at the University of Pretoria that has recently commenced to further investigate the conditions required for undrained failure to occur. The research includes laboratory and field testing to replicate these conditions in the laboratory and relate them to those found in an active tailings dam.

Jun 2, 2022

By L. Makua, K. Langa, Ndlovu. S., C. Saguru

"Global demand for the three platinum group metals (PGMs) Pt, Pd, and Rh has been steadily growing due to their widespread use in emission control, industrial catalysis, and numerous other applications. The primary supply of PGMs is currently the mining industry. However, legislative, operational, and environmental challenges faced by many mining operations result in erratic supply patterns and deficits for these three metals. Naturally, recycling scrap material to recover Pt, Pd, and Rh has therefore aroused research interest in the scientific and engineering communities. Given that the majority of the global demand for these three metals is driven by emission control devices, autocatalytic converters are therefore a significant source of already processed and beneficiated PGMs for recycling.Although pyrometallurgical processes dominate the recycling industry, hydrometallurgical processes may offer certain advantages, chief among these being lower emissions and reduced energy consumption. In this investigation, scrap converter material was leached in acidic chloride media and the leach liquor subjected to solvent extraction and cloud point extraction (CPE) to recover the PGMs. The recoveries for Pt and Pd using Alamine 308® dissolved in kerosene with decanol as a modifier were 95% and 81%, respectively. The maximum recoveries using CPE after complexation with 2-mercaptobenzothiazole (2-MBT) in the presence of Triton X-100 and tin (II) chloride dihydrate as a reductant were 97% for Pd, 96% for Pt, and 91% for Rh. This demonstrates that CPE, although relatively new, may have some potential as a breakthrough PGM recovery technology. IntroductionPlatinum group metals (PGMs) have found a wide variety of applications in different industries due to their chemical, catalytic, and physical properties. PGMs are essential in automobile emission abatement, petroleum refining, and the pharmaceutical and electrical industries (Carabias-Martinez et al., 2000). However, the abundance of PGMs in the Earth’s crust is low and their primary sources are fast depleting (Crundwell et al., 2011).The high demand for PGMs, together with social, economic, environmental, and political aspects related to their production, has led to concerns about their future supply. This has resulted in renewed interest in recycling end-of-life material. PGM supply from recycling has doubled over the last decade, and currently contributes 22% to global supply (Johnson Matthey, 2016). This is in part due to the environmental, social, and operational challenges associated with mining of PGMs. For example, labour unrest, energy cuts, declining commodity prices, and declining ore grades have collectively made the mining landscape economically challenging (KPMG, 2015). On the other hand, the high content of PGMs in secondary sources makes their recovery from the increasing quantities of waste products important from a resource conservation viewpoint. Moreover, PGMs obtained through recycling have been demonstrated to have a 1% environmental cost from a life-cycle perspective, whereas PGMs obtained from mining contribute the remaining 99% of environmental cost (Bossi and Gediga, 2017). Sustainable resource recovery and the economic incentive of recycling have therefore been major drivers to justify investment in PGM recycling. The autocatalytic converter industry accounts for about 70% of the global consumption of these metals (Johnson Matthey, 2016) and therefore recycling spent converters has become a major source of Pt, Pd, and Rh (Figure 1)."

May 1, 2019

By J. A. L. Napier

It has been established that significant time-dependent stope convergence may occur over time periods of hours and days in certain hard-rock gold and platinum mines. The source of this time-dependent behaviour appears to be associated with both pre-existing discontinuities and with mining-induced fractures that form near the stope face. These induced fractures may be associated with blasting processes and may also be formed in response to high stress concentrations in the unmined regions immediately ahead of the stope face. In shallower platinum mining operations, time-dependent behaviour is, however, observed to be much less marked unless some form of specific mining-induced fracturing occurs. One particular case of considerable interest is the time-dependent behaviour that is found to be associated with the formation and deployment of crush pillars. The purpose of the paper is to present a simple limit-equilibrium computational model of this behaviour that is sensitive to both the formation sequence and the size of planned crush pillars in a mine layout. This model provides a useful means to optimize the sizing of crush pillars, and at the same time may be used to identify potentially hazardous circumstances in which pillars may not crush in a stable manner.

Jan 1, 2012

By A. F. Silva, J. M. G. Sotomayor, V. F. N. Torres

Geotechnical monitoring plays an important role in the detection of operational safety issues in the slopes of open pits. Currently, monitoring companies offer several solutions involving robust technologies that boast highly reliable data and the ability to control risky conditions. The monitoring data must be processed and analysed so as to allow the results to be used for several purposes, thereby providing information that can be used to manage operational actions and optimize mining plans or engineering projects. In this work we analysed monitoring data (pore pressure and displacement) and its correlation with the tension and displacement of the mass of an established failure slope calculated using the finite element method. To optimize the back-analysis, a Python language routine was developed using input data (point coordinates, parameter matrix, and critical section) to use software with the rock mass parameters (cohesion, friction angle, Young’s modulus, and Poisson’s ratio). For the back-analysis, the Mohr-Coulomb criterion was applied with the shear strength reduction technique to obtain the strength reduction factor. The results were consistent with both the measured displacements and the maximum deformation contours, revealing the possible failure mechanism, allowing the strength parameters to be calibrated according to the slope failure conditions, and providing information about the contribution of each variable (parameter) to the slope failure in the study area.

Oct 1, 2021

The two methods for calibrating the parameters K and (alpha) for use in Gy’s equation for the Fundamental Sampling Error (FSE)–, Duplicate Sampling Analysis (DSA) and Segregation Free Analysis (SFA)–, are described in detail. A case study using identical broken reef material from a Witwatersrand-type orebody was calibrated using the DSA and SFA methods and the results compared. Classically, the form of Gy’s equation for the FSE raises the nominal size of fragments given by dN to the power of 3. A later modification of Gy’s formula raises dN to the power (alpha), the latter term being calibrated with the coefficient K in the DSA and SFA methods. The preferred value of for low-grade gold ores used by sampling practitioners in the mining industry is 1.5. A review of calibration experiments for low-grade gold ores using the DSA and SFA methods has produced values of K that vary between 70 and 170 and values of in the range 0.97 to 1.30. The average value for is shown to be 1, rather than 3 as originally proposed in the classic form of Gy’s equation or the industry-preferred 1.5. It is suggested that for low-grade gold-bearing ores the equation for the FSE should raise dN to a power of 1. Such an equation for the variance of the FSE greatly simplifies the characterization of gold ores, now requiring only the calibration of K for a given mass and established fragment size. The implications of the simplified equation for the heterogeneity test are that, provided the fragments have been screened to within a narrow size range, any particular size will return a value for K that is acceptable for use in the sampling nomogram.

Jan 1, 2017

By G. Song, S. Yang

"SynopsisThe development of techniques and equipment for single-pass high-seam longwall mining in China is reviewed. Some methods used to obtain the fractured zone height are discussed. The boundary between the caved and fractured zones is generally not clear, but one important difference that distinguishes the two zones is that horizontal compressional forces exist only between blocks of fractured zone. Based on this, a theoretical method is presented to investigate the destabilizing modes of the main roof (by sliding or rotation) and is used to determine if the main roof is in the caved or fractured zone. This method considers the mining height, immediate roof thickness, bulking factor, main roof thickness, main roof strength, main roof periodic weighting interval, and vertical stress in the overlying strata. To verify the method, a representative physical model of Wangzhuang coal mine is developed. The movement of overlying strata and fractured zone height are thus obtained. According to the results, the total collapsed height of the strata reaches about 70 m above the coal floor; the first main roof bending interval is 45 m and periodic bending distance is 10–15 m; the maximum strata subsidence is around 62 mm; and the presence of a three-hinged arch causes the fluctuation of subsidence in the same stratum level. Based on the theoretical and experimental analysis, the middle of the main roof is considered the boundary between caved and fractured zone; the caved and fractured zone heights are about 21 and 49 m respectively.IntroductionThick coal seams (>3.5 m) account for about 44% of the coal rescource/reserves and more than 40% of the total annual production in China (Zhao 2004; Wang 2009). Three mining methods – slice mining, top coal caving, and single-pass high-seam longwall mining – are frequently used in underground thick-seam operations. Compared with slice mining and top coal caving, high-seam single-pass longwall mining significantly simplifies the mining processes, reduces roadway excavation, and improves coal production, productivity, and recovery. The favourable economics, high production, and technical simplicity make high-seam single-pass longwall mining one of the most popular underground mining methods for thick coal seams in China."

Jan 1, 2015

By L. M. Tavares, H. A. Petit, P. P. Cavalcanti

Iron ore pellets experience degradation during handling due low energy impacts and stresses. These impacts lead to volume or surface breakage of the pellets, creating fine particles that accumulate along the process. Some of the particles are in contact with the air and are dragged as dust into the workplace or into the atmosphere. The aerodynamic size of the particles allows them to be potentially inhaled by the workers or to be carried to nearby communities. Dust emission becomes critical at specific locations, such as transfer chutes and stockpiles. This work models the dust generation that occurs in an iron ore pellet stockyard using a methodology that combines experiments to quantify the degradation of the pellets to mathematical modelling of the dust emission mechanisms in the stockyard. First, laboratory tests were used to characterise the amount of mass and the particle size distribution of the dust generated by degradation under controlled conditions. Second, the flow over a stockpile of an iron ore pellet company was simulated with the aid of computational fluid dynamics (CFD) simulations. Third, semiempirical equations linked the data generated by simulations and laboratory tests to estimate the dust emission in the stockyard. Finally, results were validated with measured field data, obtained with a novel online measurement tool. Results suggest that the approach is able to predict dust emission with sufficient accuracy and can be used for further implementation in the management of operations and decision making in order to reduce the environmental impact of the process. Keywords: Iron ore pellets, abrasion, dust emission, mathematical modelling

Jan 1, 2020