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By G. Ferrara, P. Bozzato, B. Chinè

To obtain a reliable comparison between conical and cylindrical separatory vessels used in dynamic dense-medium separation systems, the following two studies were made: a study of the swirling flow in conical and cylindrical vessels using laser-Doppler velocimetry (LDV) and a study of the behavior of suspensions of magnetite and ferrosilicon in water in both separators. It was found that the tangential velocity profiles differ in conical and cylindrical vessels. In a conical cyclone, free vortex conditions affect the largest part of the vessel, while in a cylindrical vessel, the motion approaches that of a forced vortex, i.e., that of a centrifuge. It was also found that the conical vessel manifests higher centrifigal acceleration and very strong densifying (the density differential) and, hence, the density gradient inside the vessel is large. It was found that the cylindrical vessel manifests lower centrifugal acceleration and low densifying, i.e., low-density differentials, and hence, the density gradient inside the vessel is small. Also, due to greater densifying, the density cut-point in cyclones is higher than the operating medium density, while, in cylindrical vessels, the density cut- point is closer to the medium density. The results provide an interesting description of the differing behaviors of conical and cylindrical vessels, and they help explain the differences in pegormance between the two types of separators.

Jan 1, 2000

By C. J. Pritchard, R. E. Gundersen

I commend the authors on examining an interesting method of shaft lining and attempting to improve safety in a dangerous environment by remotely installing the shaft lining. It is important, though, to not trade off shaft lining safety for long-term mine safety by lining an intake shaft with combustible material. In many operations, the mine utilities are installed next to the lining of the intake shaft, supplying a possible ignition source if damaged or overloaded. Although the polyurethane studied in the paper is of relatively high density, no mention is made of the flame retardency of the samples. When exposed to heat, small amounts of polyurethane will melt and bum but are extinguished when the heat source is removed. If large surface areas of combustible polyurethane - a tunnel or mine shaft- are exposed to ignition heat sources that are not removed or extinguished early, an uncontrollable, fuel rich fire may result with high carbon monoxide levels, low oxygen concentrations, and greatly increased burning rates that can have devastating effects on the ventilation system and miners' safety. Although these heavy polyurethanes lend themselves to a "pouring" application around the annulus of the shaft liners and is a continuous process, past experience with atomizing and frothing polyurethane spray systems indicates a high degree of maintenance and operator attention to the nozzle1 delivery system. This may require more miner's time to be spent in the shaft maintaining the foam system, negating some of the safety advantages of the system. If polyurethane foams are to be used in mines, they must be of the highest quality available in terms of flame spread, heat emitted while curing, and contaminants emitted while curing and spraying. In addition, employees must be given adequate training and be thoroughly informed in all aspects of polyurethane foam safety procedures. As witnessed by our fellow miners in South Africa, we can accept no less.

Jan 1, 1991

By P. K. Gbor

Over 2 million tonnes of smelter slags containing Co, Ni and Cu are produced and stockpiled in Canada each year. Smelters also produce SO2 an air pollutant. Recycling these materials will maximize the value of natural resources and minimize possible environmental impacts of industrial operations. The objectives of this research were to understand the characteristics of slag, the reaction of slag with SO2(aq) and to develop a novel method for extracting Co, Ni and Cu from slag using SO2. The properties of slags from different smelters were studied, with emphases on the form of Co, Ni, Cu and Fe. Leaching experiments were carried out in a batch reactor. The parameters investigated included concentration of SO2(aq), stirring speed, particle size, temperature and pH. O2 was introduced into the reactor to study its effect on the system. The amount of each of Co, Ni and Cu in all the slags was less than 1%, while Fe was between 30% and 45%. The form of Co and Fe was primarily oxide, however, significant portions of Ni and Cu were present as sulphides. SO2(aq) alone, was effective in leaching Co and Fe. Cu dissolution was negligible due to precipitation. The presence of O2 improved Cu and Ni extraction. SO2(aq) was found to be more effective in leaching slag than H2SO4 at the same pH. The leaching process was modelled using the shrinking core model. A scheme was proposed for the extraction of Co, Ni and Cu from slag using SO2(aq). The process makes efficient use of SO2, separates Co and Cu at an early stage, and minimizes the consumption of base during Fe separation.

Jan 1, 2003

By Kostas Fytas

A computer package has been developed which is capable of evaluating alternative production schedules (long and short range) of an open pit mine once the ultimate pit limits have been defined. The model user can choose the best alternative according to cash flow considerations and the particular mining conditions. The package involves a combination of simulation for long range production scheduling and linear programming for short range scheduling. The simulation model based on an iterative technique, maximizes the profits within each long scheduling period (eg. year or sixth month period), and at the same time defers stripping as long as possible. A minimum exposed ore quantity is kept at the beginning of each mining period. The final long range schedule satisfies a set of constraints for each period, mainly ore quality and quantity constraints for ore and waste. An interactive smoothing procedure at the end of each period results in a schedule that is feasible in terms of mining practice. Finally a cash flow analysis model carries out cash flow calculations so that the user can select the best alternative. The linear programming model, based on a certain number of shovel regions, defined interactively by the model user, calculates the optimum production schedule for each short period under consideration, so that a set of constraints is satisfied.

Jan 1, 1986

By Orville L. Lantto, Bruce K. Cantrell, Keith S. Olson

Approximately 10 to 20 pct of exhaust aerosol mass from diesel equipped vehicles operating without exhaust controls is attributable to particles in the 1.0-10 µm size range. Recent in-mine measurements suggest that, in addition to reducing overall exhaust aerosol mass concentrations, the use of waterbath exhaust conditioners (water scrubbers) may significantly reduce the concentration of diesel exhaust particles larger than approximately 0.5 µm aerodynamic diameter. This implies that a measurement system that discriminates between diesel and mineral dust aerosols on the basis of aerosol size would have different accuracies depending on use of a water scrubber. To test this hypothesis, laboratory measurements were made by the Bureau of Mines to determine diesel exhaust aerosol penetration efficiency through a water scrubber as a function of size and engine operating conditions. Size distributions were determined using a Micro-Orifice, Uniform Deposit Impactor. Concentrations of CO, CO2, NO x, O2 and gas-phase hydrocarbon concentrations were also measured. No significant effect was found for these gases other than a slight decrease in NOx emissions. The water scrubber served as a moderate aerosol control with reductions in mass concentrations of about 29±9 pct under most conditions. This reduction occurred for particles less than 0.2 µm and greater than 0.6 µm aerodynamic diameter. Coal dust aerosols are mostly larger than 0.8 µm aerodynamic diameter. With use of a water scrubber, separating the aerosol into two parts, greater and less than about 0.8 µm aerodynamic diameter during sampling, results in collection of diesel and coal dust aerosol accurate to within 10 pct. Without a water scrubber in operation, such a sampling protocol would only be accurate to approximately 15 pct.

Jan 1, 1991

By C. H. Lee, F. Concha, L. G. Austin, P. T. Luckie

The Concha-Almendra batch-settling treatment for concentrated suspensions of single-diameter spheres has been modified. This modified version is used in the differential equations of simple batch settling of a particle size range of different density materials. The equations have been programmed for a finite difference solution, and a number of test cases have been analyzed. By cutting the settling column at a fixed fractional height, the material below this height can be treated as underflow and the material above as overflow. The predicted partition curves obtained from treating the settling in this way have Sharpness Indices in the range of 0.5 to 0.6, in the absence of dispersion effects. The Sharpness Indices decrease as dispersion increases. When the feed size distributions are of the Rosin-Rammler form, the partition curves (corrected for the bypass fraction) fit the Rosin-Rammler equation with reasonable accuracy. The predicted Sharpness Index of a component of a mixture of two materials of different densities was unchanged from that of the same materials settled alone. However, for the same, feed size distribution and volume fraction of solids in the feed, the heavier material settled faster, and the lighter settled slower, than when settled alone.

Jan 1, 1993

By Y. Ye, J. D. Miller, E. Pacquet, S. Gopalakrishnan

From laboratory research activities during the past five years, it is now evident that effective flotation separations can be achieved with the high-capacity air-sparged hydrocyclone. The effectiveness of the equipment has been demonstrated on a pilot-scale basis (1 tph) for a number of mineral commodities including coal, copper, gold, and industrial minerals using a nominal 5-cm x 50-cm system with separation efficiencies equivalent to those achieved by conventional flotation machines. The high specific capacity of the air-sparged hydrocyclone (100 to 600 tpd per cubic foot of cell volume for a nominal 5 cm x 50 cm system compared to 1 to 2 tpd per cubic foot of cell volume for conventional columns and mechanical machines) makes it especially well-suited for the pro- cessing of low-grade ore. As a result of these pilot-scale successes, a number of plant-scale test programs are planned for larger-diameter units to establish wear characteristics and scale-up relationships for design purposes. The status of the air-sparged hydrocyclone technology is reviewed in the present paper, including a description of the system, fundamental considerations of flotation mechanisms, general discussion regarding the effect of control variables, and the significance of feed characteristics.

Jan 1, 1988

By Dragomir D. Kocic

At its lowest hierarchical level, the mine ventilation control system consists of local systems controlling the concentration of noxious gases in the air currents of various branches of the ventilation network, and of local systems controlling the air flow rates through those branches in which it is not necessary to regulate the concentration of any noxious gases. With respect to the noxious gas concentrations, the manipulated variables are air flow rates, while with respect to the air flow rates - the manipulated variables are aerodynamic resistances of the corresponding branches. Since there is an interdependence between the air flow rates in the ventilation network branches controlled through changes in aerodynamic resistance, the total system of mine ventilation control falls in the class of multivariable systems. The treatment of individual branches of a controlled ventilation network as separate plants presupposes an absolute autonomy of local systems, which is an untenable concept. The present paper is an attempt to break up the multivariable control system of the noxious gas concentrations and the air flow rates into autonomous (mutually unrelated) local control systems. The plant discussed is a ventilation network of arbitrary configuration with a single main axial fan. Because of the complexity of the problems involved, the multivariable control system of noxious gas concentrations and air flow rates has been described as linear, with constant and lumped parameters.

Jan 1, 1980

By James Sheridan

Risk can generally be defined as the possibility or chance of a loss corresponding to uncertainty. Analysis is the process for examining the components and relationships of a complex system in order to determine the impacts of its variables. Risk analysis is therefore a course of action taken to analyze multiple variables of a system that is subject to speculative outcome. When risk analysis is applied to mine development, it becomes the means for assessing and measuring those parameters that contribute to economic viability. Mining can be a risky business. It is composed of an assortment of factors that contribute to its financial outcome. Of these factors, the most consequential output is the net present value (NPV). Applying risk analysis in mine development can identify those factors having the most significant impacts and applying limits to those impacts. Mathematical computer modeling using Monte Carlo techniques is available and can be customized to suit specific mine development ventures. Essentially, those parameters that are not fixed values can have input variability. Rationally deduced engineering estimates should be the basis for determining the initial values applied as inputs in the models. Risk analysis therefore provides the user a means to identify, rationalize and assess the risks in mine development in an informed, calculated, and unbiased format for making economic decisions.

Jan 1, 2002

By C. B. Manula

The total systems model for underground coal mine planning and design (Master Design Simulator) is a set of computer programs which encompasses a realistic reproduction of the interaction among the geological, environmental, equipment, methods, and the health and safety aspects of the mining plan (Manula, Rivell, and Ramani, 1975). The total system includes nine major subsystems: geology and reserves, methane generator, roof support, subsidence, production, rail haulage, cost, water generator, and ventilation (Fig. 1). The subsystems themselves are divided into two categories based on the frequency with which they are executed: (1) event-oriented subsystems are executed either independently or at discrete intervals when certain specific events occur; and (2) dynamic subsystems function interactively during the production cycle, transferring data to or requesting data from event-oriented subsystems only when specific events occur (Fig. 2). Each subsystem has been tested and validated independently (Ramani, White, and Sutton, 1974; Ramani and Owili-Eger, 1974; Manula, Mozumdar, and Jeng, 1974a; Manula, Mozumdar, and Jeng, 1974b; Manula, et al., 1974; Manula, Rivell, and Sanford, 1974; Ramani, McClay, and Burgos, 1975; Manula and Owili-Eger, 1975; and Ramani, Owili-Eger, and Manula, 1975).

Jan 1, 1979

By Ching L. Wu, Bolek Chytrowski, Peter Frobenius, Dale Roberts

During the past few years, the first stage of the Waste Isolation Pilot Plant, (WIPP) which consists of an exploratory shaft, a ventilation shaft, shaft stations, access drifts and test rooms, has been under construction. Construction of the first stage is scheduled to be completed in the summer of 1983. The U.S. Department of Energy (DOE) is constructing the WIPP facility to demonstrate the feasibility of disposing of defense-related nuclear waste in the bedded salt deposits of the Delaware Basin near Carlsbad, N.M. The project promises to be a proving ground for underground nuclear waste storage facility design and construction. The completed underground plant will extend over an area of 647,500 m2 (160 acres), and 1,360,500 metric tons (1,500,000 tons) of salt will be excavated. The first stage of the project, the Site and Preliminary Design Validation (SPDV) Program, provides for full-scale construction of the major components of the underground facility. Shaft construction and underground development are described. Rock bolt installation is performed as required for roof control. Geotechnical instrumentation is installed to monitor the behavior of underground openings. Direct observation of the constructed shafts, shaft stations, and underground openings is expected to confirm the adequacy of the underground design and allow full-scale construction of the first U.S. underground nuclear waste storage facility to proceed.

Jan 1, 1983

By Todor P. Stefanov

The thermal draught within ventilation networks of underground mines depends to a great extent on the shafts depths and degree of change of ventilating air density. When the topology and space orientation of a ventilation network are constant, the dynamics of thermal draught is determined mainly by the variations of the heat- and mass-transfer processes in the openings. This conclusion follows from the well known basic theoretical notions about the thermal draught-aerostatics and thermodynamics [1]. Quantitative evaluation about the influence of heat- and mass-transfer processes on dynamics of thermal draught could be made through: direct mine measurements, special experiments and with the aid of mathematical models. The first two methods produce sufficiently accurative results, however it is very difficult to perform such studies, particularly when these processes are evaluated within a large scope and multifactor aspects. Mathematical models of thermal draught allow to overcome easily these difficulties. The degree of accuracy of results, obtained by this method, depends to a great extent on the degree of approximation achieved by the model to the real heat- and mass-transfer processes. In this paper, with the aid of Bulgarian developed method, algorithm and computer program V WT-4 - for complex evaluation of ventilation parameters in active mine ventilation networks [2], the authors investigate some changes in the thermal draught. Basically, the method is an iterating evaluation of an open set of equations, which represents relationships between the following basic parameters of a mine ventilation network: [ ]

Jan 1, 1980

By Eckart Schneider

PROJECT The Walgautunnel is the heart of a new hydro scheme in Western Austria, which is supposed to go into operation in 1984. With it's length of 21,4 km, a bored diameter of 6,25 m and an internal operating pressure of 13,5 bar it will be the largest water conducting tunnel in Austria. Even compared with some of the country's world-famous highway tunnels like 'Tauern' or 'Arlberg' it has to be regarded as a major tunneling project. Owner is the Vorarlberger Illwerke AG, a power corpo¬ration founded in 1924. In the course of the past 6o years this company developed the 'Werksgruppe Obere Ill - Lünersee', consisting of 6 major power plants and 4 large reservoirs. The installed capacity is 1.114 MW in the generating and 529 MW in the pumping mode. The total power output amounts to 1.916 GWh. The water for power generation comes from the glaciers of the Silvrettagroup, as this corner of the Austrian Alps between Vorarlberg , Tyrol and Switzerland is called. The Walgauwerk utilises all the water accumulated by the existing schemes as well as the so far unused run-off water, passing by in the river Ill. The intake to the Walgau-Tunnel is situated right beside the lower reservoir of the recently completed pump-storage plant 'Rodund II' at an elevation of 64o m above sea level. Though the head is a mere 162 m, 1 billion m3 of water will generate 356 GWh per annum. The power-output of the two Francis turbines is 86 MW.

Jan 1, 1983

By Steven A. Tesdahl

Maintenance means keeping equipment running or restoring it to operating condition. However, the 21st century will usher in a broader need for equipment management, a cradle-to-grave strategy to preserve equipment functions, avoid the consequences of failure and ensure the productive capacity of equipment. Profitable future mining operations will have reduced the 35% of operating costs typically spent on maintenance and the unfavorable impact of downtime that often multiplied these costs by 300%. They will survive those operations that tried to carry outdated "maintenance" thinking beyond the year 2000. These surviving operations will have applied modern management techniques, technology and information to align the efforts of people with the needs of equipment. Future managers will use equipment management as an integral part of an overall production strategy. They will see it as a means of raising productivity, improving performance and minimizing downtime to maximize profitability. This global thinking will characterize all operations planning to survive in the competitive world of mining of the next century.

Jan 1, 1998

By Louis W. Cope

Placer material, regardless of the value of gold, tin, titanium, diamonds or other value mineral it contains, remains worthless until the valuable constituent is separated from the bulk of the material. This separation is no mean task when considering that a gold placer containing 0.01 to 0.03 oz/yd3 of gold, which is the range of grade of typical placer deposits; yes, 1:3,000,000! It is a miracle that gold placer mining has ever been profitable. Added to the situation of minute quantities of gold are several other problems, such as the gold often being very finely divided, the association of boulders and clay, and running and/or ground water. As each placer deposit is individual and distinctive, many factors have to be considered in planning the method for gold recovery from the alluvial deposit. Previous chapters have shown that in the exploration and sampling stage, some knowledge of the amount of each size of coarse and fine material which occurs in the deposit as well as the size distribution of the discrete gold particles is determined. Unless informed to the contrary, as a general rule of thumb, plant designers will consider that 40 percent of the material will pass through a 5/8-inch screen or punched plate hole. Also that there may be an occasional nugget and normally considerable amounts of fine gold. The designer must

Jan 1, 1992

By Andrew B. Cecala, Robert J. Timko, Christopher J. Pritchard

As mine venti1ation systems become more complex, it is increasingly difficult to supply fresh air to the working face. A high percentage of the fresh air entering a mine never reaches the working face because it short circuits through stoppings and doors, exiting the mine without being utilized. When ventilation limits are reached, the only option presently available to most U.S. mines is to sink a new shaft - an extremely expensive undertaking that many companies cannot afford. Reusing or recirculating a percentage of ventilating airflow may make it possible for some mines to remain operational while maintaining a safe environment for workers. Such a study was performed at a trona mine in Granger. Wyoming. A vane axial fan was used to recirculate approximately 9.4 m /s (20,000 cfm), 25% of the total ventilating airflow, back to the working face area of a three-entry continuous miner section. This study involved two evaluations: the first when the face was 305m (1,000 ft) from the recirculation fan, the second when the face was 610m (2,000 ft) from the fan. The evaluation consisted of monitoring the mine atmosphere for respirable dust, methane (CH4). carbon monoxide (CO), and nitric oxide (NO), at various locations in the intake and return entries. Several shifts were monitored, with and without the recirculation system in operation. With the recirculation system in operation. there was an increase in measured respirable dust concentrations in the intake along with a slight increase in the mine temperature (1.67°C). There were also some minor changes in CO and NO levels in the intake entry, but this seemed to depend on the proximity of the instruments to the turn-around point for the diesel haulage vehicles, than to whether the recirculation system was or was not in use.

Jan 1, 1989

By Robert F. Randolph, Arnold C. Love

This study describes the positioning strategies used by operators of radio controlled underground mining machines. The observations have implications for task design to insure adequate visibility and task performance. Radio controlled continuous coal mining machines have allowed many mines to increase safety and production by placing human operators away from the hazards of an unsupported mine roof. The operator is no longer confined to an operators component and visibility can be enhanced with proper positioning. Also, the operator is no longer subjected to the shocks and vibrations of the mining machine. The operator's ability to move freely about the worksite requires a careful analysis of several competing safety and production considerations. Thirty-five operators of remotely controlled continuous mining machines were observed using work sampling followed by interviews to determine where the operator was positioned and what he or she was watching. These observations were separated according to the task being performed by the continuous mining machine and summarized as position density plots. The specific task of tramming in reverse through a crosscut is analyzed along with more general summaries of the time miners spend outside of the policy-defined operating positions while performing other production tasks. The operators were: observed to frequently vary from company standard operating procedures (SOP). Many of the variations from company SOPS were based on task demands. Interview responses indicated that portions of the mining task place visibility demands on the operators that are difficult to accommodate within existing procedures. An improved task design process would include these visibility requirements. Also, the operator's vision can be augmented thorough video cameras in future teleoperated mining systems, and the operator can be increasingly isolated from environmental hazards.

Jan 1, 1992

By E. Bane Kroeger

Plane failure in rock and soil slopes is usually caused by the intersection of a discontinuity plane with a slope face. Normally, these discontinuities are weaker than the surrounding intact rock and provide a surface along which movement can occur. The discontinuities are typically analyzed to determine the stability of a particular slope. One problem that arises when evaluating the stability of a planar feature is that the method for planar analysis is two-dimensional and has difficulty dealing with planes that are oblique to the slope face. Several researchers have stated in the past that planar failures are a special case of wedge failure. Wedge failures occur when two planar discontinuities intersect in an unfavorable manner. Sliding can occur along the intersection of the planes. Wedge failures are very common in rock slopes and can often be problematic. Wedge stability analysis methods are typically three-dimensional and can analyze planar features that are very oblique to the slope face. Wedge failure analysis can also be used to determine the stability of planar features with good success. One advantage to using wedge analysis over plane analysis is that a tension crack can be both non-vertical and oblique to the slope face in the wedge analysis. This allows the engineer to more closely match field conditions. It also negates the need for the engineer to have two separate software packages to analyze a single slope, saving both time and money.

Jan 1, 2002

By P. J. Macbeth, R. F. Overmyer, K. K. Nielson

INTRODUCTION Over the past several years, Ford, Bacon & Davis Utah Inc. (FB&DU) has been involved in research, development, and engineering activities concerning uranium mill tailings; specifically of interest in this paper are efforts in radon gas exhalation and surface flux attenuation measurements, environmental influences on radon fluxes, and basic research into radon gas behavior. Our projects have included measurements of effectiveness of specific cover material in attenuating radon surface fluxes from mill tailings, characterization of the parameters in cover materials and tailings that influence radon exhalation, long-term tailings stabilization planning, determination of the effects of moisture on radon emanation and diffusion from ores, and determinations of baseline radon fluxes. Our efforts have been sponsored by U.S. Department of Energy, the U.S. Nuclear Regulatory Commission, Tennessee Valley Authority, the commercial uranium industry, and we are presently engaged in efforts for the U.S. Bureau of Mines. Much of the material presented in this paper is taken from our published reports on radon flux attenuation and tailings stabilization (Macbeth et al., 1978; Nielson et al., 1980; Rogers et_ al., 1980; Thamer et al al., 1980). CHARACTERIZATION OF RADON GAS EXHALATION One dimensional diffusion theory based on Fick's laws is the basis for the modeling of radon gas migration through large quantities of porous tailings and cover materials applied over the large areas of tailings piles. Fick's first law relates the gaseous concentration gradients to the flux over the entire surface and defines the effective diffusion coefficient De as: [ ]

Jan 1, 1980

By Michael J. Shulman

During-all phases of exploration limited resources must be allocated among exploration methods and among exploration prospects. This problem,, an exercise in optimization, can be treated mathematically. For three objective functions - maximizing return, return to cost and return minus weighted cost - theoretical results exist which enable the exploration planner not only to gain valuable insight into the problem, but also to estimate the resources necessary and to allocate these resources for' at least-two stages of exploration prior to initial field work. A two-stage exploration strategy for volcanogenic-sedimentary iron in northern Egypt illustrates, how the results can be used. Stage 1 consists of an aeromagnetic survey to map the structure of five areas;" stage 2 consists of field reconnaissance work. For the three objectives, lower bounds are calculated as a function of exploration expenditures for each stage in order to facilitate the choice of an-optimal exploration budget. Nomograms are then presented which, based on the Kuhn-Tucker Conditions', enable the incremental allocation of airborne and reconnaissance effort among the five areas.

Jan 1, 1989