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By D. Nilsson, B. Aaro

To find out if there are any potential savings in "in-the-wall haulage for open pits," the Swedish companies ASEA and Kiruna-Truck in 1984 gave us financial support to study this solution in more detail. In the study, the use of in-the-wall haulage was studied for three hypothetical open pits with different sizes and shapes. Annual ore production rates also varied. A part of this thor¬ough study was published by the Swedish Mining Research Foundation (see Report F 8444, "The Use of Electric Trucks and In-the-Wall Haulage in Open-Pit Mining," in Swedish with an English Summary). Although in-the-wall haulage is of interest in some open pits, the local terrain is important for profitability. The authors do not think in-the-wall haulage is of any major interest for the mining industry. The following is a summary of some of the author's findings. To place the haul road in the wall is not of interest in open pits with declining metal contents in the bedrock. In such mines, the volumes from the haul roads will not yield reve¬nues. The cost per m3 for an underground haul road is much higher than for a haul road in the pit. This means that it is only of interest to place the haul road in the wall when mining gets deeper than about 50 m (165 ft). In-the-wall haul roads will reduce flexibility in the pit, and it will make necessary the use of smaller equipment with lower productivity and higher costs per ton. As Hustrulid, et al. show (Table 4), excavation savings of material hauled are very low, $0.10 to $0.25/t (0.09 to 0.27 per st). The extra operating cost, due to lower productive equip¬ment, will normally be much higher and thus destroy the whole idea with in-the-wall haulage. If electric trucks are profitable for in-the-wall haulage, it is normally also profitable to use electric trolley assist for trucks on the haul road in the pit. But the profitability of using electric power is different in different countries, and depends on the relation between the cost for electric power and diesel fuel. In the US, diesel fuel is inexpensive compared with electricity, but in a country like Sweden, diesel fuel is much more expensive. In most open pits, the trucks have to move a considerable distange from the loader until they reach a final haul road in the wall. A trolley line along such temporary haul roads will be exposed to flyrock. It is normally less expensive to perform rock support from the open pit than from underground ramps in the wall. In Fig. 8, Hustrulid et al. gives the impression that the underground haul roads will be very close to each other. This is seldom the case. Figure 1 shows a haul road in the wall. The haul road passes through each cross-section only once. Ar¬ranging a reliable dewatering and rock support system from only one underground ramp is probably impossible. Many more drifts are therefore necessary. In our report, we also studied what would happen if the final pit slope is increased by 5°, using underground drifts in the wall. Our conclusion was that the extra cost for drifts, rock support, pumps, etc. destroys the whole idea, and that it was better to accept a higher stripping ratio with the haul road in the pit and to use conventional low-cost open pit equipment. Finally, we think it is would be difficult to try to reduce the safety factor when determining the slope angle by moving the haul road in the wall. Minimizing the risk for pit wall collapse is also important with in-the-wall haulage, primarily because men and equipment will be working in the pit, but also to guarantee many accesses between the pit and the haul road in the wall.

Jan 1, 1989

By S. Majumder

The longwall face supports are provided to prevent the caving of the immediate roof at the working face. The thickmess of the immediate roof may be obtained from the working height and the average bulking factor. A strong bed, capable of withstanding maximum bending moment among the beds within the above roof, can be identified. The load on the strong bed may be obtained by finding out the overlying beds that will rest on the former during sagging. The limiting span, L, for this unsupported loaded bed, as a beam with built-in ends, can be calculated. The support density required at the supported span (face to goaf edge), a, during the first fall, is now estimated with a view to have a safety factor 2 at the face end, when bending failure occurs at the barrier end. However, the above value is inadequate to provide necessary safety margin against bending failure at the face end, during the subsequent cavings as cantilever, for L/a less than 20.6. d For the latter case a separate expression is derived.

Jan 1, 1986

By J. F. T. Agapito, C. E. Brechtel, C. J. Gilbert, T. R. Kelly, R. C. Karlson

A multi-lift benching cut-and-fill mining method was adopted at Asamera Minerals (U. S.), Inc.'s Cannon Mine to balance the goals of high productivity and near 100 percent recovery with the need to minimize subsidence. High strength backfill was required to provide effective long-term overburden support and safe working conditions during the pillar recovery operations. Operational designs were evaluated using numerical modeling to predict global mine stability and backfill strength requirements. These models incorporated sequential mining of primary stopes, backfilling of primary stopes and recovery of pillars, and have been transferred to the mine operations group for use in evaluating different production sequences. The modeling has provided a better understanding of the structural response of the mine to excavation of the primary stopes and removal of the pillars. Predictions about the transfer of overburden weight to the abutments, pillars and cemented fill pillars have provided a criterion to evaluate mining plans and production sequence. Comparison of these predictions with actual mine structural performance has developed confidence in the design of the mining method and in the geomechanical design methods being employed. Primary mining of the B-North zone is nearly complete, and the stopes have been generally stable, as projected in the original design. Some exceptions have occurred where locally stope pillars have failed due to the presence of adversely oriented sheared beds. Recovery of the pillars between the

Jan 1, 1987

By Carmine Nappi

This article examines some important trends in East-West trade in non-fuel mineral products, in particular the recent decline in Eastern bloc exports of such products. It relates this decline to an improvement in Soviet terms of trade; a deterioration of geological and locational conditions in Soviet mineral production; an increased integration of the CMEA-countries based on a regional pooling of resources; and a surge in demand for mineral commodities in Eastern countries coupled with a slowdown in Western economic growth. Finally, it evaluates to what extent these determinants may evolve in a different direction during the eighties and thus foster East-West mineral trade. Eastern1 bloc trade in mineral products is a subject of perennial interest in the West. When exports increase, there is much talk of the dangers of dependence on the East for critical fuels and materials. When imports increase, talk turns to the danger of Soviet dependence on Western supplies. A common fear is that the USSR may not only seek to meet its own resource needs but also deny the access of Western countries to these resources. This paper, like three other recent additions to the lite5ature on Soviet foreign trade of commodities2, tries to put matters in clearer and cooler perspective. The first section evaluates the structure of East-West mineral trade and tries to identify some important trends of recent years. The second section relies on supply and demand factors, as well on the Comecon connection (the intra-Council for Mutual Economic Assistance mineral cooperation) to explain why there has been a reversal in the mineral trade trend since the end of the seventies. The last section evaluates the factors which could stimulate East-West mineral trade in the future.

Jan 1, 1986

By K. R. Weller, U. J. Sterns, P. S. B. Stewart

Complex sulphide ores usually require very fine grinding of the whole ore stream to achieve adequate liberation. Recent improvements to DSM screens have made this device appear competitive with cyclones in circuits treating these ores. A DSM screen and cyclones were tested in pilot-scale grinding of a copper, lead, zinc ore and the resultant data were used to establish a multi-component simulation model. The model was used to assess the influences of mill environment, circuit flows and pulp densities, final product properties (size, relative size and SG of the values, and liberation characteristics) and a second grinding circuit on capacity and product size distributions, and hence define conditions for which one device is better than the other.

Jan 1, 1992

By Steven J. Page, Robert A. Jankowski, John A. Organiscak

Quartz dust is one of the most significant ongoing health concerns in coal mining today. Since initial verification of this health risk in coal mines by the Mine Safety and Health Administration (MSHA) during the early 1980’s, the Bureau of Mines has conducted numerous studies in underground mines to identify the sources of quartz dust and its fundamental characteristics. The most significant underground quartz dust source was found to be the continuous miner excavating rock. The percentage of quartz was also found to be higher in the smaller size fractions of the dust when significant amounts of rock was mined by the continuous miner. Furthermore, the quartz content in the continuous miner return remained fairly consistent at greater distances from the face. Another important factor speculated to affect quartz dust generation and its size characteristics is the morphology of the quartz bearing rock. The quartz content of dust between mines had notable differences with respect to the differences between kaolinite content in the dust. Low kaolinite content was associated with higher quartz content and increased size segregation of quartz into smaller dust particles, indicating that the free silica bonding matrix in the rock is most likely another factor responsible for quartz dust produced through mechanical comminution.

Jan 1, 1991

By Robert Dimock

The successful management of large dispersed mineral operations is dependent upon accurate and timely decision making. The plant locations providing the data to support the decision making process are generally separated by distances which elongate the time required to accumulate the data at a central focus. The need exists to examine operational data in context with total mining operations, and in relation to target projections and historical trends. The use of microcomputers for local information tasks is an established capability. Grouping microcomputers into local-area networks is rapidly becoming established. Interconnecting multiple local-area networks to provide a free flowing stream of information between all operational locations of a mining enterprise is possible with use of some emerging technologies. The benefit of the potential interconnections and rapid transfer of information can readily be seen from a production reporting perspective. Numerous other advantages exist' by providing routing to enable a workgroup to be established between separated work functions.

Jan 1, 1989

By Kumar Vikram

Rich reserve of prime quality limestone in east khasi hills of India and abundant quantity of low-cost fuel (Natural Gas) in Bangladesh made Lafarge Surma cement limited a joint venture company of Lafarge France and Cementos Molins Spain to envisage an integrated cement plant at Chatak District-Sylhet) a bordering town of Bangladesh. Since April 2006 the single flight 17 km Long Belt Conveyor (LBC) built by M/S Larsen & Toubro Ltd. (L&T), ECC Division an engineering and construction company of India, is feeding limestone/shale to the Lafarge Surma cement plant. This engineering marvel was created on a turnkey basis with technology support of M/S Aumund France

Jan 1, 2008

By David M. Spatz

Anyone who scans the modern literature will realize that concepts of leadership have changed during the past few years. And like change in so many areas of our culture, change has been rapid in the perception of leadership. Many traditional leadership traits are still highly regarded and many of the longstanding personal qualities that improve leadership abilities continue to be effective. In the past, though, leaders could rely on dominance, intimidation and fear to accomplish their goals through the actions of others. In contrast, effective leaders today build relationships, provide vision and guidance, and strive for communication, cooperation, trust and employee satisfaction. Earnest leadership involves moral issues like honesty, integrity, fairness, empathy and respect. Modern leaders guide, facilitate and coach, rather than command and direct. But how well equipped are managers in the mining industry to meet the leadership challenges that lie ahead? More importantly, is he or she becoming a better leader? This article outlines steps to becoming a more effective leader. Management consultants stress that the changes that lie ahead will demand swift organizational adjustments and managers who are equipped to lead during times of change (Boyett and Conn, 1991; Hesselbein et al., 1996; White et al., 1996 and Kotter, 1997). Becoming a better leader begins with self-assessment and a commitment to unify all who depend on one's leadership abilities. Simply defined, leadership is leading others. Contrary to common practice, leadership is not about advancing one's own career but about bringing out the best in others. The beauty of modern leadership, however, is that the path of serving others creates synergies. They benefit everyone, including the leader, those the leader manages, upper management and company shareholders.

Jan 1, 1998

By Chen Bing Ching

Hsikwangshan mine is located at Lengshuijiang city of Hunan Province. It is a stibnite deposit. The orebody is gently inclined; its ore run is about 3 km in length and extends about 1.2 km along the direction of dip. The thickness of the ore body is non-uniform and varies from 1.5 to 20 m. The orebody is mined by room and pillar and cementation filling. The nine levels at the ore top have been mined out. At present, the mining operations are at the eleventh level. The mine ventilation system consist of three main fans which are installed in different shafts and are ventilated in parallel. The main fans can be connected to each other through comprehensive return air drifts. Two of the fans are located underground and the other on the surface. The general configuration of the ventilation system is shown in Figure 1. Because the operating areas are scattered and vary frequently as well, it is difficult to bring about air flow control. If the air flow cannot be varied with the variation of production in time, the air quantities are not sufficient in some areas and are excessive in others. However, it is also a problem to coordinate the three main fans with one another to obtain the optimum economic cost of operation. In order to make the air delivery fit the variation of production continuously to satisfy the ventilation requirement of the working faces and to reduce power consumption, we have made a great effort over five years to install an automatic control system for mine ventilation at the Hsikawangshan mine. The key part of this system is a microcomputer. The whole system consists of a console, which is located at the surface; the circuit network that is provided for signal transmission; air doors and ventilators at the working face; the scale air doors which are installed at the main branch airways; main fans and their speed regulating devices, as well as the sensors used for observing pressure, air quantities, temperature and the concentration of carbon monoxide, etc. This system can continuously telemeter the air quantities and carbon monoxide concentration in the main branch airways, air pressure, air quantities, electric current, the temperature of the bearings of the main fans, etc. According to the production air requirements, the system controls the ventilators or air doors which are located at the working faces and regulates the air quantity in the main branch airways. It can also choose the optimum mode of operation automatically, and regulate the rotating speed of the main fans to guarantee the least total power consumption while satisfying the total air quantity requirements. If abnormality of the observed results is discovered, the audible and flashing light alarm signal starts up automatically. The operating mode of ventilation equipment also feeds back to the control room and displays on the console. The configuration of control the system is shown in Figure 2.

Jan 1, 1987

By Patrick A. Ley

Maintaining, managing and manipulating large quantities of geologic data can be a very costly and time-consuming process. The use of computer-based methods can provide some relief to the problems of editing, displaying, and interpreting spatial geological data. ENCORE is a system designed specifically for coal orebody modeling and reserve esitmation where the geologic data is stored on a large mainframe computer and is accessed from a local work station for editing, manipulating, and producing various reports and high-quality engineering maps. The ENCORE system was designed for use by the geologist or mining engineer with little or no computer background and, additionally, it was designed to provide the geologist or mining engineer enough flexibility to be able to select the appropriate methodology for the orebody being analyzed. The various analysis tools incorporated into ENCORE can significantly increase the professional's productivity without burdening he or she with the need to understand the details of the internal computer language.

Jan 1, 1982

By D. Nilsson, B. Aaro

To find out if there are any potential savings in "in-the-wall haulage for open pits," the Swedish companies ASEA and Kiruna-Truck in 1984 gave us financial support to study this solution in more detail. In the study, the use of in-the-wall haulage was studied for three hypothetical open pits with different sizes and shapes. Annual ore production rates also varied. A part of this thorough study was published by the Swedish Mining Research Foundation (see Report F 8444, "The Use of Electric Trucks and In-the-Wall Haulage in Open-Pit Mining," in Swedish with an English Summary). Although in-the-wall haulage is of interest in some open pits, the local terrain is important for profitability. The authors do not think in-the-wall haulage is of any major interest for the mining industry. The following is a summary of some of the author's findings. To place the haul road in the wall is not of interest in open pits with declining metal contents in the bedrock. In such mines, the volumes from the haul roads will not yield revenues. The cost per m3 for an underground haul road is much higher than for a haul road in the pit. This means that it is only of interest to place the haul road in the wall when mining gets deeper than about 50 m (165 ft). In-the-wall haul roads will reduce flexibility in the pit, and it will make necessary the use of smaller equipment with lower productivity and higher costs per ton. As Hustrulid, et al. show (Table 4), excavation savings of material hauled are very low, $0.10 to $0.25/t (0.09 to 0.27 per st). The extra operating cost, due to lower productive equipment, will normally be much higher and thus destroy the whole idea with in-the-wall haulage. If electric trucks are profitable for in-the-wall haulage, it is normally also profitable to use electric trolley assist for trucks on the haul road in the pit. But the profitability of using electric power is different in different countries, and depends on the relation between the cost for electric power and diesel fuel. In the US, diesel fuel is inexpensive compared with electricity, but in a country like Sweden, diesel fuel is much more expensive. In most open pits, the trucks have to move a considerable distange from the loader until they reach a final haul road in the wall. A trolley line along such temporary haul roads will be exposed to flyrock. It is normally less expensive to perform rock support from the open pit than from underground ramps in the wall. In Fig. 8, Hustrulid et al. gives the impression that the underground haul roads will be very close to each other. This is seldom the case. Figure 1 shows a haul road in the wall. The haul road passes through each cross-section only once. Arranging a reliable dewatering and rock support system from only one underground ramp is probably impossible. Many more drifts are therefore necessary. In our report, we also studied what would happen if the final pit slope is increased by 5°, using underground drifts in the wall. Our conclusion was that, the extra cost for drifts, rock support, pumps, etc. destroys the whole idea, and that it was better to accept a higher stripping ratio with the haul road in the pit and to use conventional low-cost open pit equipment. Finally, we think it is would be difficult to try to reduce the safety factor when determining the slope angle by moving the haul road in the wall. Minimizing the risk for pit wall collapse is also important with in-the-wall haulage, primarily because men and equipment will be working in the pit, but also to guarantee many accesses between the pit and the haul road in the wall.

Jan 1, 1990

One of the major objectives of cost analysis is to enable the manager to predict the financial outcomes of decisions. To accomplish this, the future costs and benefits of alternative courses of action must be determined. If care- fully used, information about past financial performance can be helpful in this analysis despite the fact that, by definition, financial reporting information is historical. In many cases, both revenues and costs need to be estimated for decision- making. For the production manager, additional revenue is a result of increased throughput or recovery of valuable material from the stream of mined material. A reduction in operating cost, although similar to a revenue increase in its financial impact, is estimated using slightly different techniques. Generally, the estimation of cost is more difficult than the estimation of revenue. Most cost estimates have two components: an estimate of fixed costs and an estimate of variable costs. As discussed in chapter 5, the total cost of an activity during a specified period can be expressed as: Total cost = (Variable cost per unit of activity) x (Units of activity) + Fixed cost or TC = VX+F

Jan 1, 1999

By T. T. Chen, J. L. Jambor, J. E. Dutrizac

A reconnaissance study of the neutralization potential of 30 diverse minerals, 28 of which are silicates and aluminosilicates, has shown that most of these minerals contribute minimally to the neutralization potential (NP) in standard static tests. Feldspars in particular give low NP values and would not be expected to attenuate the acidity to any significant extent in acid rock drainage settings. The BET-determined surface areas of the minerals vary considerably, and the large surface areas of some of the pulverized samples have clearly affected the NP values even though the standard static-test protocol was used for the preparation of the samples. The low NP values of most silicates and aluminosilicates are in general agreement with the laboratory-determined slow dissolution rates for these minerals. The results suggest that, in the prediction of whether rocks will be acid-producing, the most realistic measure is that of the carbonate content. In mine wastes from metalliferous ore deposits, the contribution to the NP by the non-carbonate gangue assemblage is unlikely to be more than a token amount until well after the establishment of an acid rock drainage situation.

Jan 1, 2000

Conventional logging tools such as density, neutron, and acoustic have ranges of investigation of tens of centimeters. The borehole gravimeter, however, has a range up to hundreds of meters. For example, from the center of a cylindrical ore body 0.20 g/cm3 denser than its surroundings, one can detect the boundary of the ore body at a radius as large as five times its height. More complex structures can be investigated with computer programs. We can predict the usefulness of proposed gravity surveys, or decide which suggested structure is most likely to have caused the results of existing surveys.

Jan 1, 1988

By D. E. Martin, K. J. Hamill, H. J. Miller

1. INTRODUCTION In April 1979, the U. S. Nuclear Regulatory Commission issued the Draft Generic Environmental Impact Statement (GEIS) on Uranium Milling (NUREG-0511) for public comment. While the GEIS is a broad environmental impact assessment of the U.S. uranium milling industry as a whole, its primary focus is on the issue of uranium mill tailings disposal and long term isolation. In August of 1979, proposed regulations (U.S. Federal Register Notice 44FR50012), incorporating both conclusions of the GEIS and provisions of recent U.S. legislation on uranium mill tailings (Uranium Mill Tailings Radiation Control Act of 1978--UMTRCA), were also issued for public comment. This paper (1) briefly reviews the major conclusions of the draft GEIS, (2) summarizes the major elements and rationale of the associated, proposed regulations, and (3) reviews some of the major issues raised during the public comment period and changes which the NRC staff expects will be made in finalizing the GEIS and regulations. The risk in summarizing a document of such broad scope and complexity as the GEIS is that the depth of evaluation and the full technical bases for conclusions reached may not be adequately conveyed. Recognizing this, those Sections of the draft GEIS where a more complete discussion of the issues is presented are parenthetically noted in this paper in order to facilitate location of additional background and supporting material. 2. GENERAL CONCLUSIONS In the GEIS, a full range of tailings disposal alternatives was considered. These alternatives ranged from the past practice of doing little or nothing to control tailings, to advanced treatment alternatives such as nitric acid leaching to increase removal of radioactive elements and fixation of tailings in cement or asphalt. Our broad conclusions were: -Any alternative which results in the need for ongoing, active maintenance to assure long term containment, is unacceptable. (12.3.2.1) -Advanced treatment alternatives are unwarranted since they offer uncertain incremental benefits yet involve costs which may be as much as an order of magnitude greater than costs associated with acceptable alternatives. In addition, the advanced treatment

Jan 1, 1980

By R. Bruno

The problem of testing different production strategies has been studied in relation to reserves selection and a computer programme for ore production management has been developed. The case study presented concerns an underground rock salt mine in Sicily with different product lines and variable ore characteristics. A geostatistical block model of the deposit was constructed on the basis of the physical and chemical parameters that ultimately determine the type of end-product in operations according to different selection strategies were simulated using the model and taking account of both technical and 'commercial constraints. The structure of an ore production management programme running on a PC to select both the mining unit and the end-use of the ore is described.

Jan 1, 1989

By William G. Davenport, Eric H. Partelpoeg

Direct smelting of copper in a single vessel has long been the extractive metallurgist's dream. Such a process offers the important advantages of (i) maximum use of the energy available from oxidizing the sulphur and iron of the concentrate; (ii) restriction of SO2 emissions to a single constant-strength source; (iii) minimum materials handling and (iv) elimination of at least one process step. Extensive developmental work was carried out in the 1960's and 1970's to implement such single-step processes. Some of the processes reached the prototype stage and it appeared that copper smelting technology was on the threshold of a new, efficient, minimum-pollution era. Since that time, however, the forward thrust has slowed and single-step smelting is still a process of the future. The paper describes in detail the potential advantages of single-step smelting; its history and current status; and the reasons why it has not been adopted industrially. It concludes that the potential benefits of single-step smelting are powerful and that many of the difficulties encountered in the 1970's will be overcome in the 1980's.

Jan 1, 1985

By C. E. McFarland, N. W. Kirshenbaum

Cortez Gold Mines is located about 75 miles southwest of Elko, NV. Activity in the mining district dates from 1862. The current operation developed from a 1963 joint venture of Webb Resources, The Bunker Hill Company, Vernon F. Taylor, Jr., and American Exploration & Mining Co. Three years later, the USGS discovered gold on the property following a regional geochemical survey. Drilling in 1966 and 1967 identified 3.4 million st of oxide ore grading 0.28 oz Au/st, and a 1700 st/day countercurrent decantation mill was commissioned in 1969. Heap leaching started in 1971 on low grade ore and continues to the present. The mill was shut down in 1976 following exhaustion of known oxide reserves. Mobile equipment was sold and the proceeds committed to local exploration. In 1980, a new plan of operations was developed starting with existing low grade and newly discovered mill ore. The mill resumed operations in 1981 and has been operating to date. A series of metallurgical innovations, culminating in the commissioning of a new fluid bed roaster in the summer of 1990, has enabled treatment of a series of increasingly refractory ores. Current partners in the venture are Kennecott Corporation, Vernon F. Taylor, Jr., and Placer Dome U.S. Inc., manager of the operation. Idaho Mining Corp. holds a 7.5% carried interest.

Jan 1, 1992

By S. G. Hardcastle, Akbar Sheikh

In Canadian underground uranium mines, bacterial leaching is being increasingly applied to low-grade ore reserves. At Denison Mines Limited uranium operations, old stopes are used as leaching vessels; these contain a leaching muck pile brought down by drilling and blasting from the back which contained a low-grade reef. The bacteria used require atmospheric oxygen to oxidize the pyritic ore. A study was initiated to evaluate the ventilation, including oxygen supply, specifically, of flood leaching stopes during their rest period. The method of ventilation investigated was a single ended compressed air system which feeds airlines laid under the muckpile and carried through the inlet pressure bulkhead. The air delivery system was tested for post- blast integrity with a sulphur hexafluoride, SF6, tracer gas technique. This was injected into the intake compressed air line and collected at the stope's borehole exhaust in grab-sampling bags. The samples were analyzed with a laboratory gas chromatograph and results provided information on the stope's airflow, average residence time, overall clearance time, volume of the airpath and the airflow distribution. The borehole exhaust was monitored by anemometer and psychrometer and indicated that the psychrometric conditions in the stope were independent of the area's ambient climate. Evaluation of the exhaust's standard gaseous composition showed the oxygen consumption of the bacteria to be independent of the investigated flow range. Field derived results are compared with computer models, theoretical predictions and laboratory investigations.

Jan 1, 1987