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By A. E. Prokop, J. P. Rider

Previous research on continuous mining operations has shown that significant interactions exist between dust-control parameter application and the resulting respirable dust levels. However, simply increasing the level of the controls does not guarantee a reduction in respirable dust levels. Full-scale laboratory tests are being conducted to provide information to longwall operators that would assist them in selecting control parameters to reduce dust levels for mine-specific conditions. The interactions between face air velocity, shearer water quantity, drum water spray pressure, external water spray pressure and spray system design were evaluated in a simulated 2.13-m (7 ft) coal seam for two cutting directions. Locations around and downwind of the shearer were monitored to evaluate relative changes in respirable dust levels as a function of each control parameter.

Jan 1, 2003

By M. R. Fletcher

Slags generated by the ferrous metals industry can he used to produce high-value ceramic materials. The properties of these ceramics can be controlled by judicious choices of composition, additives, and processing parameters. Standard methods of materials handling, such as comminution and mixing, are as applicable to slag as to any other mineralogic raw material. In this study, foundry slag was ground, homogenized, melted, and, formed to produce a glass, and then heat treated to crystallize it into a high-strength glass-ceramic. The product shows promise as a low-cost, wear-resistant tile material that (-an he custom cast. This and similar materials made from slags and natural rocks are candidates for use as construction materials, acid-resistant piping, shock-resistant sheathing, rot-proof pilings, and, when foamed, even as pontoons.

Jan 1, 1990

By J. W. Hawkins

Understanding aquifer properties in surface mine backfill is essential for managing the groundwater flow system as well as controlling the formation and discharge of acid urine drainage. Slug tests (injection and withdrawal) performed on wells in mine spoil may yield two distinctly different response characteristics. One type of response is indicative of a porous media system. The second type of response indicates the presence of a pseudokarst system. The presence of voids or conduits in the backfill is indicated by an almost instantaneous partial recovery of the water level displacement. This rapid recovery is turbulent flow and cannot be analyzed by conventional (Darcian) methods. The turbulent flow may cause the calculated hydraulic conductivity to be less than the actual values. The slug dissipation surface area created by the voids contiguous to the test well add to the known dissipation surface area of the screened interval. This can cause the calculated hydraulic conductivity to be greater than the actual hydraulic conductivity or the aquifer to appear more permeable. Whenever analyzing data from aquifer tests in mine spoil, these potential problems must be kept in mind. In spite of the problems encountered during slug testing of spoil, the aquifer hydraulic parameters derived and the slug responses observed remain useful in the hydrologic characterization of a mine site.

Jan 1, 1994

By L. Tondo, W. Lim

The high pressure grinding rolls (HPGR) is a relatively new technology which offers the potential benefits in energy savings, circuit simplicity and improved product characteristics in minerals processing applications. Over 300 machines are now operating world-wide, but most of these are used for grinding cement clinker, and relatively few are used in conventional mineral processing. One of the rea- sons for its current limited use in the minerals processing industry is that design, scale-up and flowsheet assessment procedures are seen to be relatively immature and untested. This has reduced confidence in the potential application of the technology. To help rectify this situation an industry funded research project was undertaken, one of whose aims was to develop and prove a mathematical model of the HPGR and a scale-up procedure. Data were collected from tests using laboratory, pilot and in one case a full scale unit (2.2m diameter). A range of feed types were tested including quartz, limestone, kimberlite, bauxite, iron ore and gold ore. These data were then used to develop models of the size reduction performance, power draw, throughput and operating gap for a wide range of pressures, feed sizes, and roll speeds. This paper describes the models that have been initially developed and illustrates their accuracy using some of these data.

Jan 1, 1998

By Michael Gauna, William J. Gray, Joseph C. Zelanko

Mechanized mining in the underground coal industry has continued to evolve since the introduction of the remote control mining machine. Remote control continuous mining machines without an onboard operator’s compartment are now the norm in the industry. Extended cuts in excess of 30 ft (9.1 m) are commonplace. Unfortunately, machine operator positioning has become an often unrecognized safety concern, particularly when remote control continuous mining machines excavate a turn (turn a crosscut). From January 2001 through July 2004, remote control continuous mining machine operators or helpers have been involved in ten roof fall fatalities and eight machinery crushing fatalities. The machinery crushing incidents occurred when machines were being repositioned or moved to another place. Of the ten roof fall fatalities, three occurred while a crosscut was being turned and a fourth fatality occurred while mining was being conducted adjacent to the crosscut. The operating time associated with turning crosscuts is a relatively small part of the mining cycle. In recent years, however, a disproportionate number of fatal incidents appear to have been associated with this activity. Procedures for reducing the incidence of operator fatalities while turning crosscuts are discussed. These techniques include: minimizing the number of turn cuts, limiting turn cut depth, establishing no work zones and using visual warning devices to designate them, installing more roof/rib support in the vicinity of the turn cut, using notch cuts prior to initiating a full turn cut, and positioning the remote control miner operator in nontraditional locations.

Jan 1, 2005

By J. T. Sitarski, R. C. Emmett, L. W. Luinstra, R. P. Klepper

A mining operation in Zimbabwe produces a pyritic gold concentrate that requires oxidative pretreatment prior to cyanidation in order to liberate 80 percent of the gold values. Bio- oxidation offered the potential of increased gold recovery and was considered to be an economically advantageous alternative to the existing practice of roasting the concentrate. Based on bench- scale test work, a small bio-oxidation plant, capable of processing approximately 50 percent of the concentrator production, was installed at the mine to evaluate this approach. The plant included two 57 m3 bioreactors in series and the ancillary equipment needed to prepare the feed concentrate, to wash the leached residue prior to cyanidation, and to neutralize the waste liquor. The plant was operated for a year-and-a-half, and the conditions needed to achieve a 90 percent liberation of gold values were determined. A number of critical technical problems were encountered, and many were resolved on-site. Others have been addressed since the pilot plant was moth-balled in June 1988. Valuable insight was gained from the operation of the pilot plant, which has provided useful guidance for continued process and equipment improvement.

Jan 1, 1990

By Gordon S. Thomas

The problem of how best to exploit a reserve has vexed mine planners and design engineers since mining began. The cyclic nature of the mine design-scheduling problem, combined with the lack of good mathematical tools for tackling it, have ensured that it has been, and will continue to be, a major challenge for operations researchers working in mining. This paper reviews the current state of the art in pit optimisation and scheduling algorithms. Starting from the inception of the first block models in the late 1950's, it progresses through the developments and applications of the Moving Cone algorithm, the Lerchs Grossmann 3D Graph Theory algorithm and some of the heuristics, to the new artificial intelligence techniques that are being applied today. Along the way, it discusses the limitations of the various approaches and lays the foundations for the research that is currently being undertaken by CSIRO and others in the field of mine production scheduling and control.

Jan 1, 1996

By M. Polat, G. Ahuja, S. Chander, M. S. Klima

A study was conducted to investigate the use of flotation in combination with dense-medium cycloning utilizing ultrafine (-10 um) magnetite, as an integrated process for high-efficiency cleaning of fine coal. The objective was to produce a high-quality (low-ash, low-sulfur) coal, while reducing the amount of non-magnetic material that would need to be removed by magnetic separation. Three dense-medium cyclone tests were performed using different underflow diameters to produce feed material for the flotation separations The use of different underflow configurations caused a variation in the pulp and solids splits, which changed the clean-coal yield and the magnetite-to-coal ratio of the cyclone products, and hence, the composition of the flotation feed. Flotation was used to separate the clean coal and to recover the magnetite contained in the cyclone overflow stream. Any ultrafine refuse material, which may have reported to the cyclone overflow stream, was also removed in this step, further upgrading the product. In addition, flotation was used to improve the overall yield by recovering a middlings fraction from the cyclone underflow stream. It was found that flotation was very effective in separating the coal (or middlings) from the ultrafine magnetite, even under variable feed conditions. Magnetite recoveries approaching 99% were possible using rougher-cleaner separations in a laboratory flotation cell. This paper presents the results from this study.

Jan 1, 1995

By B. Rao Pothini, Julio H. Thaler, W. L. Finlay

The White Pine room-and-pillar mine started in 1955 from a suboutcrop, down gentle dips, and with a competent sandstone roof. As mining depths increased, extraction decreased 1% every 35 feet. Moreover, roofs became shaly and less competent, and regions of high fault and joint densities, sometimes with and some- times without high lateral stresses, increased the tendencies towards roof falls. Instinctive reactions towards narrower openings did not alleviate the problem. Careful consideration of rock mechanics principles culminated in the development of two variants to the conventional room-and-pillar mining system based on the concept of stress relief through load deflection. The two systems-EquiAdvance Cave, and PressureArch Panel Mining-are complementary. The objective of the former system is to provide relief from lateral stresses through advance caving and in the process also to achieve higher extraction at depth. Where surface subsidence cannot be allowed, the alternative PressureArch Panel concept is utilized to provide stress relief and moderately high extraction ratio. Both systems have been under trial for over three years and, together, produced over three million tons of copper ore. These two systems, and their production test results to date, are described.

Jan 1, 1977

By F. K. Letowski

Leaching of zinc sulphide concentrate by nitric acid which is simultaneously regenerated by oxygen in the reactive froth has been investigated. The reactive froth was formed by a recycled zinc sulphate electrolyte containing nitric acid, a mixture of gaseous 02 and NOx, entrapped in the foam and particles of the zinc concentrate suspended in the foam. In the reactive froth the multiple regeneration of nitric acid proceeded owing to the encapsulated oxygen and amount of nitric acid used can be considerably lower than in the continuous liquid phase. 93 to 97% of zinc was extracted to a solution containing 30 to 409/1 HNO, during 40 to 60 min of leaching in a batch reactor without regeneration of nitric acid. 50 to 100% of the consumed nitric acid was regenerated during continuous leaching in froth carried out in the ‘flotation column’-type reactor at 80 to 90°C and under pressure of 100 to 200kPa. The rate of HNO, regeneration in the froth was proportional to oxygen pressure and inversely proportional to nitric acid concentration in the froth. A two-stage process flowsheet is proposed. In the first, the reduction stage, removal of nitric acid from the leach solution directed to standard purification and electrolysis is carried out by its reduction to nitric oxide using fresh concentrate in the froth. In the second, the oxidation stage, nitric acid is continuously regenerated by oxygen in the froth.

Jan 1, 1995

By Marvin Barnes

The Sacaton mine is a relatively new open¬pit copper mine owned by Asarco, Inc., which has been in production only a few years. The mineral deposits consist of two ore zones; a western, near-surface zone minable by open-pit methods, and a deeper eastern zone which will have to be mined underground, probably by block caving. The copper deposits are related to small quartz monzonite intrusives implaced into Precambrian granites. Both rock types act as hosts for the copper mineralization, which has been secondarily enriched sufficiently to constitute ore. In the western zone the enriched supergene chalcocite has been subjected to postenrichment oxidation processes with partial destruction of the chalcocite blanket. The deeper eastern chalcocite zone has not been as severely affected by later supergene processes. Several mineralogical populations are represented in the total copper deposit. Leached capping overlies the chalcocite enriched zone, and in the western ore body there is a very irregular interface between the leached and enriched zone, which is complicated by oxide copper minerals irregularly distributed. Primary lower-grade chalcopyrite mineralization occurs below the enriched zone. When mining commenced near the top of the enriched zone, considerable difficulty was encountered in correlating the actual grade of ore in place with the estimated grade of the mining blocks. The mineralogical complexity and fairly wide spacing of diamond drilling made reliable grade estimation of individual ore blocks a difficult procedure.

Jan 1, 1979

The stability of mine roofs in permafrost zones is determined mainly by temperature and ice content. The deformation resulting from creep and temperature-dependent physical properties of frozen rock mass is the dominant component of total deformation. Roof bolts can consolidate the roof and control its deformation. In this paper, a finite-element model, for the analysis of the effects of roof bolting in a frozen rock mass is presented. The model includes the variables normally encountered when analyzing the effects of roof bolting in rocks. It also accounts for properties unique to frozen rock mass: creep deformation and the change in Young's modulus with changing temperature. The model shows that, as bolt spacing decreases, there is a trade-off between increased roof stability from increased mechanical support and decreased stability from increased conduction of heat into the frozen roof. This result was expected and shows that, for a given set of rock-mass conditions, including ice content, temperature, particle size, etc., there is an optimum bolt length and bolt density for minimum roof deflection.

Jan 1, 1997

By Jakub Nalaskowski, Vamsi K. Paruchuri, Marcin Niewiadomski, Jan D. Miller

Spherical oxide particles can be used as mineral fillers, paint pigments, polishing materials and catalysts. They can be prepared using various techniques involving precipitation and spray drying. Melting of particles in a high-temperature flame is a simple technique which produces particles with a high degree of sphericity and high smoothness from non-spherical feed material. In current research the preparation of alumina particles by melting the feed material in a hydrogen- oxygen flame is described. The spherical particles obtained were characterized using scanning electron microscopy, atomic force microscopy, energy dispersive spectroscopy, X-ray difiaction, Fourier transform infrared spectroscopy and electrokinetic measurements. The particles pre- pared exhibit a high degree of sphericity and a low level of roughness. The chemical composition of the material after preparation is unchanged, however the crystallographic structure is different. It is suggested that alumina changes from pure a-alumina to a mixture of different transitional forms. Changes in the electrokinetic behavior were also observed for alumina which can be attributed to the strong dehydration of the oxide surface and/or change in the crystallographic structure. Also, the possible accommodation of trace amount of impurities in the surface zone during melting cannot be ignored.

Jan 1, 2003

By G. B. Misra

Attempts have been made during the last three decades to establish a correlation between respirable dust dosage and the degree of progression of pneumoconiosis or silicosis amongst mine workers. Fair correlation has been possible only after the Johannesburg Pneumoconiosis Conference established mass concentration of dust in the respirable size range as the true measure of the incidence of pneumoconiosis. The author had established a good correlation between cumulative total respirable dust dosage as well as cumulative silica dosage and the category of pneumoconiosis according to ILO classification at Mufulira Copper Mine. Studies at Hutti also show a linear relationship between the average cumulative dust and silica dosage md the category of pneumoconiosis. However the maximum tolerable cumulative dust dosage for - prolonged exposure is 182 mg month/m3 - at Hutti as compared to 275 mg month/m3 at Mufulira. Similarly the maximum tolerable silica dosage for - prolonged exposure is Only 35 mg month/m3 at Hutti as compared to 55 mg month/m3 at Mufulira. The TLV for dust concentration at Hutti works out at 0.6 mg/m3 compared to 1.1 mg/m3 at Mufulira and the TLV of silica concentration at Hutti is 0.117 mg/m3 compared to 0.25 mg/m3 at Mufulira. The Hutti figure is however, closer to the TLV of 0.1 mg/m 3 recommended by ACGIH.

Jan 1, 1987

By R. V. Ramani, K. V. K. Prasad

Forecasting mined grade variability from exploration information is a challenging task in mine planning. For effective planning of activities, such as workplace scheduling, ore grade quality control, etc., development of realistic numerical simulation models of in situ variability from exploration information is an essential first step. Geostatistical conditional simulation (GCS) is one of the well known methods to develop such simulations. Recently, stochastic fractal models based on the assumptions of scale-invariance and self-similarity have been used for developing simulations of several natural phenomena. It has been suggested that these models might permit the generation of short-scale variability information from exploration data with greater fidelity when the data satisfies the attendant assumptions. In this paper, stochastic fractal models are reviewed, and their relationship to geostatistical variogram models is examined. A comparison of the two simulation techniques has been performed using coal sulfur data from a mined-out coal property. The results of this comparative analysis are also presented in the paper.

Jan 1, 1993

By B. R. Reddy, P. V. R. Bhaskara Sarma

A process for the separation and recovery of cobalt and nickel was optimized for a sulfate leach solution of Indian Ocean sea nodules containing 80 g/m3 cobalt and 730 g/m3 nickel using the sodium salt of Cyanex 272 as the extractant. Extraction of both cobalt and nickel increased with an increase in equilibrium pH and extractant concentration. The extracted species with respect to cobalt appears to be CoA2.3HA. More than 99.6% extraction of cobalt was achieved using 0.04M solvent in a three-stage counter current simulation at the A:O phase ratio of 3:2. The co-extraction of nickel was about 0.6% only. Two-stage countercurrent stripping of cobalt from the loaded organic at an A:O ratio of 1: 7 resulted in a pregnant electrolyte enriched by 0.8 kg/m3, thus enabling a ten times overall enrichment of cobalt. From the cobalt free raffinate, nickel extraction was carried out using 0.25M solvent in two countercurrent stages, and the extraction showed a 99.8% recovery. Stripping of nickel from the loaded solvent was carried out at an equal phase ratio with the advance electrolyte of pH 1.75 in two countercurrent stages. The nickel stripping efficiency was found to be 98.8%. The present studies indicated that complete recovery of cobalt and nickel was possible using nine counter-current stages.

Jan 1, 2002

By F. Calizaya, I. Duckworth

An ore-pass is an underground conduit used to transfer ore from one level of a mine to another. This method of transportation depends on the elevation energy only; therefore, it has very low operating cost. However, two inherent problems may hamper this process: (1) hang-ups of interlocked pieces and (2) a sudden release of trapped air from the ore-pass (belching). These events can create unsafe conditions with the potential for catastrophic consequences. This paper presents a summary of CFD modeling studies that were formulated to investigate the pressure build-up and belching of dust and rock particles from long ore-passes. The models were calibrated against field data collected from an actual ore-pass system and used to predict the airflow-pressure distributions in the mine for various ore feed rates and ore-pass geometries. P.T. Freeport Indonesia’s Grasberg ore-pass system (240,000 tpd capacity) was used for this purpose. This system comprises of four active ore-passes, with each being more than 600 m in vertical length.

Jan 1, 2004

By R. Karl Zipf, Z. T. Bieniawski

Based on the premise that understanding the fundamental mechanics of fine fragment formation in coal will lead to reducing coal dust problems in mines, a thorough study of the mechanics of crack propagation in coal was undertaken. Two approaches were considered in an attempt to understand and quantify the fracture process as it relates to fine fragment formation. In one approach, theoretical models are used which describe the size of the fracture process zone which may relate to the amount of structural damage and fine fragment formation on the new crack surface. The other approach is based on the relationship between crack tip stress intensity and new surface appearance for mixed mode crack propagation. Preliminary observations with a scanning electron microscope support the premise that fine fragment formation is related to fracture surface appearance and fracture initiation conditions. An experimental program designed to supply the input parameters for these approaches is presently underway. The experimentation features mixed mode fracture toughness testing, using a special mixed mode testing rig, and fracture velocity measurements with a crack propagation gage. The interpretation of this data via the approaches discussed will provide further under- standing of the mechanics of fine fragment formation in coal.

Jan 1, 1986

By John M. Marek

INTRODUCTION This paper is for the mining manager, generally meaning mine superintendent, chief engineer, and others who have control over the mine operation responsible for feeding the insatiable beast called the mill or processing plant. The author is not from mine management but from the field of technical mine engineering consulting. This fact illustrates the authors gross inadequacy to comment on this subject. On the other hand, the consulting profession allows visits to many mine operations, worldwide, supplying many different commodities. The common complaints and misunderstandings that have been heard worldwide will be shared within the following few paragraphs. In order to discuss optimization, one must first define it. There certainly will be many papers in the following sessions that claim to optimize something, or other, based on a technical definition of cost, efficiency, throughput or some other quantifiable basis. DEFINE OPTIMIZATION Actual optimization of an operation is not as easily achieved as most of us in the technical world would like. Borrowing terminology from the technical world, the boundary conditions and the optimization function are usually not well defined. The problems in many cases do not lend themselves to quantification. Typical management procedures are to establish goals or strategies that set the approach or general boundary conditions to the overall plant. Three simple cases are listed below as possible corporate strategies to optimization of a mine-process plant complex. 1) Maximized discounted cash flow 2) Minimized long term cost of product 3) Maximized resource utilization This list seems simple when presented in the above context, uncluttered by constraints such as market position, capital availability, or environmental limitations. Sometimes the best approach is a prioritized combination of the above items. Typically, the Board or CEO has established these strategic priorities. They may

Jan 1, 1991

By C. Flores, R. Y. Wan, M. C. Fuerstenau, T. S. Jordan

Most oxidized manganiferous silver ores contain manganese associated with silver and are not amenable to conventional silver extraction methods, including direct standard cyanidation techniques. However, it is known that dissolution of the manganese might yield a suitable material from which silver can be recovered by conventional methods. In the segregation process, which involves heating the ore with a small quantity of salt and with a source of reducing agent to 680°C - 800°C, the silver is volatilized, migrates and is reduced to metallic silver particles. Conventional cyanide leaching is applied to the resulting calcine. Laboratory investigations have been conducted on a manganiferous silver ore analyzing 0.02 % Cu, 0.09 % Pb, 0.33 % Zn, 12.55 % Mn and 187 g/ton Ag to study the influence of roasting temperature and time, quantity of salt and quantity of coke on silver extraction by the application of segregation roasting followed by cyanidation. Cyanide leaching tests were carried out at room temperature on the segregated calcines produced by using an alkaline cyanide solution containing 2.0 lb/ton of NaCN at pH 11 for 72 hours. The results indicated that 97.8 % of silver extraction can be attained at 850°C, with 2.0 % salt, 1.0 % Coke and 30 minutes of roasting time.

Jan 1, 1993