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In 1988, at the request of members of the Society for Mining, Metallurgy, and Exploration (SME) Inc., the president of SME formed Working Party No. 79, Ore Reserve Definition. Its mission was to develop guidelines for the public reporting of exploration information, resources and reserves. These guidelines were published in the April 1991 issue of Mining Engineering,(pp. 379-384), and as a document entitled "A Guide for Reporting Exploration Information, continued on an ad-hoc basis until 1996. Working Party No. 79 was then renamed the SME Committee on Re¬sources and Reserves and became a standing committee. Since 1994, the Council of Mining and Metallurgical

Jan 1, 1999

By V. K. Karra

Research by various workers has shown compressive comminution to be significantly more efficient than impact comminution. Based on the work being done at Nordberg, wet compressive crushing in cone crushers producing finer and finer products, shows the feasibility for partial substitution of conventional tumbling mills. For further substitution, major challenges are processing of fine feed slurries in cone crushers to realize a significant size reduction of the particles, and understanding the limitation of the gyratory type units in terms of finest product achievable. The paper addresses the challenges involved, efforts at Nordberg, and the implications to conventional comminution flowsheets.

Jan 1, 1989

By Holger Witthaus

Various seam conditions had been examined during longwall production at RAG German hard coal mines. Based on the field data a prediction method was developed considering separation planes as influencing factors for roof falls at the face. The characteristics, orientation, and density of separation are important factors on the occurrence of roof falls. In addition to geological influences, the performance parameters and geometry of support must be considered and optimized for the prediction and prevention of roof falls in a longwall face. Based on detailed macro- and microscopic investigations of e coal seams considering various ranks of carbonization, the origin and characteristics of separation planes are determined and a strata-specific influence on roof fall is interpreted. It was assessed that the coal face fracturing and displacement caused by the stress depend on the floor and roof strata conditions due to activation of the existing separation planes. Field studies are shown giving an idea of influences by advance rates of the longwall face, support delay and hydraulic pressure to the roof control in longwall faces in depth of more than 3000 ft. The analysis includes the application of specific software developed for longwall roof fall prediction as a result of a European research project. The software allows a laminar graphic display of roof fall probability, which is used for recognition of risk areas and risk management.

Jan 1, 2009

By Atta Ur Rehman, Kwame Awuah-Offei

This work evaluates the effectiveness of similitude theory in predicting the draft of a rubber tire loader’s bucket during initial penetration. The work uses 1:16 and 1:8 scale models of an 18-t capacity load haul dump, scaled with the Buckingham Pi theorem, to evaluate the prediction of draft and penetration under different operating and different dynamic conditions. We hypothesized that the observed relationships between draft and longitudinal penetration and the operating and dynamic conditions will hold true for an upscaled model. Similarly, we hypothesized that the results of experiments conducted with the smaller scale model should predict the draft on an upscaled model.

By M. Sengupta

Vast reserves of oil shale occur in the Green River formation in Colorado, Utah, and Wyoming. The room-and-pillar mining method is applied to flat-lying sedimentary deposits. Important elements in mine design include mine orientation, vertical location of the mining zone, selection of roof horizons, room span and pillar size, barrier pillar size, panel dimension, and layout of mains and submains.

Jan 1, 1988

By L. Max Scott

[INTRODUCTION] The significance of gamma exposure to uranium miners has been recognized only in the last few years. Most ore deposits which have been underground mined, were 1% or less U308. Full-time mining of this grade ore can result in exposure exceeding 1 Rem per year. Several companies in Saskatchewan are planning to mine recently discovered ore bodies which contain ore pods in excess of 10% U308. The purpose of this paper is to present dose data which can be used to estimate gamma exposure from high-grade ore deposits, and to present mining techniques which will minimize miner exposure. [Dose Estimates] When one has a point source of radiation, the conventional laws of isotropic emission and the fall of dose with square of the distance, make dose calculations relatively easy. However, if one has an area source of infinite thickness, the assumption must be made that a uniform gamma flux is being emitted from all faces of the source (Figure 1). Dose from gamma emission would fall off only due to air attenuation up to such a distance at which the source can be considered to be a point source (usually 10 times the largest dimension of the source). For short distances, such as to the cab of ore-moving equipment, the dose reduction would be negligible. Assumptions: 1) Radium and gamma-emitting daughters are in secular equilibrium with uranium. 2) All gamma radiation is emitted from radium and its daughters. 3) Average gamma energy of 0.8 MeV. 4) Uranium ore has a linear attenuation coefficient of 0.212 cm-1, similar to that of Si02(sand). 5) Uranium ore density of 3.0 gm/cm3. 6) One curie of radium emits 6.84 x 1010 photon/sec(l). 7) One percent uranium ore contains 1 x 10-8 Ci/cm3 of radium. 8) Each MeV of 0.8 MeV photon results in a dose of 2.10[ ]10-3 mRem/hr. per MeV/cm2 -sec. 1) 9) Minimum ore depth 100 cm. 10) Linear attenuation coefficient instant for radium gamma of 0.374 cm -1 [ ] The photon flux from a radioactive slab is computed as follows: [Flux = photon/cm3 3(1) 2u S] [where, us is the linear absorption coefficient of s the source (uranium ore).] Thus, 1% uranium ore, in the form of an infinite slab, emits 1613 photon/cm2-sec, which results in a dose rate of 2.71 mRem/hr. Since the dose from an infinite slab is the same as a dose from an infinite radius right circular cylinder, the only correction needed from the above dose rate is for finite radius and point of dose above the end of the cylinder (Figure 2) [Dose = mRem/hr. {E2(bl) - E2(b3) - [(E2(b1sec[E] -E2(b3sec[E]]cos[O](1)]

Jan 1, 1981

By B. A. Ferguson, P. W. MacMillan

INTRODUCTION Jack Spalding in Deep Mining, Chapter 3, states: "In deep mining, to start stoping an orebody without a definite plan of operations covering the whole extraction from beginning to end is to invite serious trouble. Any stoping without plan is bound to leave, toward the end, a number of pillars, remnants, or promontories, which, as they are reduced in size, are liable to fail and cause a general collapse. The object of stope planning is so to mine an orebody that at no stage in the operation is a remnant left. Definition: `When a block of ore is stoped in such a way that eventually a small piece is left entirely surrounded by stoping, that piece is termed an island remnant.' The design of and adherence to a plan of stoping is of greater importance in preventing rock bursts than the type of ground control adopted." This, of course, is just an example of the old axiom that an ounce of prevention is worth a pound of cure. And in this context, mining does not have to be very deep before poor ground conditions, made worse by lack of good planning, force a mining company to resort to more expensive methods of ore recovery. At all times when rock pressures become excessive, stoping plans must be devised to avoid the creation of pillars or promontories. This is most readily recognized when mining deep narrow ore veins by open methods. It would be a little more general to say that stoping systems should seek to avoid or minimize localized concentrations of rock stress. Ore bodies, and particularly those at Falcon¬bridge, can reach large widths and leaving pillars for extraction at a later time cannot be avoided. However, good planning can reduce the effect of factors which create bad ground conditions. The necessary ingredients of planning are time and information: time to formulate and revise ideas; information, gained in exploration and early development, to provide the fullest knowledge of the ore body. In this way, plans become meaningful. It must always be kept in mind that the long-term results of good planning often require decisions not compatible with reaching earliest full production. PLANNING FOR DEEP MINING At Falconbridge Nickel Mines ore bodies presently being found and prepared for production lie at depths in excess of 610 m (2000 ft) below surface. Current un¬derground exploration is being directed towards favor¬able locations below 914.4 and 1219.2-m (3000 and 4000-ft) depths, at most mines. The Onaping Deep ore body has been traced to a depth of 1508.8 m (4950 ft) and at Falconbridge mine the ore body is under develop¬ment below the 5000 level. Planning for more efficient ground control has therefore become an important con¬sideration of future mining. HYDRAULIC BACK FILLING AND CUT-AND-FILL STOPING The advent of hydraulic back filling has led to the increasing use of flat-back cut-and-fill stoping methods during the last ten years, until it has probably become the commonest method of mining deep ore deposits in Canada. These methods account for almost all the pro¬duction from Falconbridge Mines at this time. The ad¬vantages claimed are: safety-a minimum of open ground; very little supplementary ground support (except rockbolted backs and walls); greater selectivity and flexibility in extraction (horizontal extension of the vein, high grade ore stringers going out into the walls); and high productivity and increased mechanization in larger ore widths. There is some difference of opinion as to whether hydraulic classified tailing back fill provides any great resistance to initial closure of the stope walls. The evi¬dence available suggests that it does not. Addition of portland cement to the back fill just prior to placement has produced a material having the properties of a weak concrete. It is confidently expected that this material will assist greatly in the efficient mining of pillars and generally in giving increased ground support to oppose closure. PRINCIPLES OF STOPE PLANNING BY LONGITUDINAL FLAT-BACK CUT-AND-FILL Again quoting from Spalding: "The practice of con¬trolling output by stopping and starting stopes is bad¬stopes once started should proceed without interruption, and control must therefore be obtained by altering the rate of stoping. It is therefore necessary that faces should normally advance at a medium rate which can be boosted, if required, to give temporarily a greater output. Modern high productivity stoping methods, however, generally aim to achieve and maintain the maximum output possible from a stope. Control of tonnage and grade, therefore, becomes a more complex problem requiring very careful scheduling and organiza¬tion. With the basic method of flat-back overhand stoping using hydraulic fill, it is not usually possible to mine in such a way as to avoid creating pillars." Consider an ore body as illustrated in Fig. 1. It is 274.3 m (900 ft) high by 228.6 m (750 ft) long on strike and averages 4.6 m (15 ft) in ore width. It rep¬resents approximately 907 000 t (1,000,000 st). The ore body could be mined by advancing one long single face through all the levels from bottom to top. A total face length of 228.6 m (750 ft) would accommo¬date five 45.7-m (150-ft) stopes. See Fig. la. Each of these stopes would produce approximately 1814 t/m (2000 stpm) and the ore body would require a period of ten years for complete extraction. A possible im¬provement to this scheme might be to stagger the indi¬vidual stope faces. See Fig. 1b. Problems of hanging wall failure can be reduced by this arrangement. From the point of view of reducing travelway maintenance, it is good practice to employ a stoping system of retreat. Stopes farthest from the shaft are mined first so that on completion, haulage drifts below can be aban

Jan 1, 1982

By J. A. Herbst, S. L. Blust

Optimization of blasting, crushing, and grinding operations is filled with challenges. One of the more difficult tasks is accurately sampling and determining the size distributions of blasted and crushed materials at a reasonable cost. The task is difficult because of the large size of fragments and the tonnage involved. However, the size distribution measurements are necessary for models that predict the performance of mine-through-mill operations. This paper is concerned with the use of video sampling for this task at National Steel Pellet Company's operation in Keewatin, Minnesota Data gathering, data analysis, model building, and mine-to-mill simulation are all described.

Jan 1, 2000

By Walter E. McCulloch

Installation of flash flotation in the copper concentrator of Freeport Indonesia in late 1986 and 1987 increased gold recovery by about 13%, while also increasing concentrator throughput. Copper recovery remained essentially the same. Flash flotation has since been included in a recently completed mill throughput expansion to 32,000 t/d (35,300 stpd) as well as in a current expansion to 52,000 t/d (57,300 stpd).

Jan 1, 1991

By V. Svalbonas

In the establishment of a useful failure criteria for complicated, fabricated structures, two basic areas of knowledge must be evaluated: 1. The accuracy of the applied stress analysis must be known in comparison with actual stresses encountered during mill operation. 2. The metallurgical fracture properties of the basic materials must be established with reference to non-destructive examination (NDE) methods to be used during fabrication. Both of the above can only be evaluated by use of full scale tests on the structures, and/or components, of interest. This paper describes such tests and comparisons as applied to several large mill installations. In addition, weldment component fatigue tests arid flawed casting strain gauge tests are also discussed. The combination of such test results allows for the formulation of failure criteria directly applicable to the fabrication and examination procedures followed in grinding mill structure manufacture.

Jan 1, 1982

By Carlos Kovacs-Figueroa

Because of air pollution control regulations in the early 701s it became necessary to clean effluent gases from the Morenci Reduction Works Smelter. Feasibility studies indicated that this could be accomplished most efficiently by using the smelter gaseous effluent to feed a sulfuric acid plant. Complying with air pollution control regulations thus created the need to dispose of relatively large amounts of concentrated sulfuric acid. Typical Morenci ores contain substantial amounts of oxidized copper minerals which are not recoverable by flotation. In addition, there is low sulfide, high oxide material which can be leached, but is not suitable for flotation concentration. Early in 1971 sufficent data had been accumulated to show that a leaching plant for the Morenci tailings could use a large portion of the sulfuric acid produced by the Smelter. It was determined that for all practical purposes leaching of the available copper was complete in two hours or less. Capital expenditures could not be justified for more than two stages of counter current decantation for recovery of the leached copper, and the plant was designed with these criteria in mind.

Jan 1, 1976

A number of factors have resulted in the widespread use of computer simulation as a day to day tool in the mineral processing industry, as is the case in most engineering disciplines. A number of simulation packages, both steady-state and dynamic are now available commercially. The developers of such software face an enormous challenge in maintaining and improving the packages due to the complexity inherent in such systems. This problem known as the software crisis faces all developers of complex software. Object-oriented (OOP) and design is a methodology which has been developed to try and minimise this problem. The use of the OOP paradigm in the devlopment of simulation software for mineral processing systems is explored in this paper and it is found that OOP techniques have major benefits for the design and development of such software.

Jan 1, 1992

By P. Legast

Serpentinite minerals are available in large quantities from asbestos mining activity. The granular or fibrous serpentite, at temperatures greater than 700°C, will recrystalize into olivine, forsterite and enstatite. These minerals have been used and studied for decades and are environmentally safe. The thermal process changes the composition and properties of the original mineral. Therefore, the substitution of the original mineral cannot simply be made and the applications for these new synthetic minerals need to be redeveloped. The two original mineral forms allow a wide range of new synthetic raw materials which can be adjusted to specific applications or custom needs.

Jan 1, 1993

By Ing Newton Amegbey, Joseph Afari Idun

Ashanti Goldfields Corporation (Gh) Ltd., the leading gold mining operation in Ghana is embarking on a major rehabilitation and expansion programme to increase gold production. A new southern ventilation district was targeted for this programme, and the ventilation system was to be reviewed in order to cope with increased production. The southern ventilation district was previously ventilated by booster fans. This arrangement was however, inadequate to solve the problem of the need to reduce smoke clearance time of the production areas, and to generally improve the underground ventilation conditions. Fan maintenance costs increased with the number of fans installed in the booster fan arrangement, and the fact that the fans were installed underground made maintenance difficult and took longer time. It became necessary to consider the installation of a new main fan on the surface. This paper gives an overview of the ventilation systems of the southern ventilation district of the mine before and after the new main fan was installed. The field tests carried out on the fan to ensure that it complied with supplier's specifications are described, and the new system analysed to see it's effectiveness.

Jan 1, 1993

By J. J. Lewis

The chapter describes how several OR techniques assisted the planning of a major new UR project. A computer system was developed to analyze reserves distribution and potential of possible underground layouts, to assist in choosing the location and size of access points. Manpower requirements of each proposed working plan were assessed and matched with the potential recruitment of skilled and unskilled labor. Computer simulations were used to determine the optimum capacity for coal preparation plant and surface bunkerage. For the overall assessment of robustness, financial modelling and risk analysis techniques were adapted to the special requirements of new mines.

Jan 1, 1979

By Harry Perry

The decline in coal production has been slowed somewhat in recent years, mostly because of the industry's own remarkable success in holding the line on costs, and partly as a result of advances achieved through industry and government research. Nevertheless, the decline has continued. This downward trend combined with the increased productivity of labor through mechanization that has been necessary for coal to retain even its present competitive position, has resulted in a serious reduction in the number of miners required and large scale economic dislocations in mining communities. For coal to regain a more favorable position it must either expand existing markets by reducing the cost of the delivered product and thereby compete more effectively with other fuels or develop new uses that will require large tonnages of coal. New markets involving several million tons, which would look extremely attractive to most other United States industries--including the organic chemical industry--would do little to alleviate the coal industry?s depressed economic conditions, While such small new markets would be welcome, an industry producing over 400 million tons per year must look to large tonnage outlets for solutions to its long range economic problems.

Jan 1, 1964

By Qingxia Liu, Chen Tingzhong, Chen Jin, Yang Xiaosheng

It is common problem (Huan Xinsheng, etc.,l992)for most iron ore concentrators in China that the concentration of thickened tailings is lower than the need for economical transportation. To solve the problem, two methods of improving the properties of the existing thickeners and using separate thickening technique have been applied in three iron ore concentrators. However, in the case of the Daye iron concentrator a new technique of cyclone-flocculated sedimentation was developed by present authors and verified by experiments. The new technique employs efficient hydrocydone designed by authors in conjunction with a low cost flocculant named KLP.

Jan 1, 1995

By Neal Cochran

I consider it a distinct privilege to be afforded the opportunity of discussing the program of the Office of Coal Research. The principal activities I will be discussing are concerned with projects that fall in the utilization area. I will, to be sure, touch on other activities of the Office of Coal Research but am not equipped to discuss these activities in great detail. Research activities of the Office of Coal Research are conducted on a contract basis with commercial, academic, and non-profit institutions. We do not operate any laboratories of our own but we do maintain a very small staff for technical and administrative supervision of our contractors. We have consistently attempted to interest outstanding, creative, and practical research organizations in our program. We are constantly seeking the imaginative new idea that will materially reduce the cost of mining coal, transporting it, or will process it into a more convenient and useable form. In assessing proposals and in developing our program, we search for the improvements that will develop large-scale new uses for coal or coal products or those that support these new uses.

Jan 1, 1963

By Toru Nakao

Fujiwara Quarry had been operated by the "Glory-hole Benchcut" mining system. But because it became difficult to meet the expanding demand by this system, "Central- shaft MCP" mining system was introduced in 1979. This MCP is equipped with a hydraulic walking system manufactured for the first time in Japan. Nowadays "Central-shaft" mining is the common method at limestone quarry in Japan. In this method crushing facilities are usually installed at the bottom of the shaft. But at this quarry, MCP was adopted and installed on the quarry surface. Many MCPs are operated in foreign countries, but in Japan few examples exist. Normally, Japanese limestone deposits lie in high mountains, which is obviously inconvenient for MCP to walk in and around the quarry. This difficulty, however, was overcome at this quarry, and MCP has been operated satisfactorily and efficiently for about 2 years after it was installed. In order to overcome the difficulty, in "Cen¬tral-shaft MCP" mining system, two mining methods are employed, the one is "SLICE-DOWN" mining method, and the other is "BLOCK-CUT" mining method.

Jan 1, 1982

By Frank F. Aplan, Edward C. Dowling, Richard R. Klimpel

Several series of time-instantaneous samples were taken on a cell-by-cell basis in the flotation circuits of the Pinto Valley, AZ copper concentrator. These samples were used to calculate mass flow rates throughout the circuit using standard, deterministic metallurgical formulae and also a variety of more statistically oriented techniques. Because of the magnitude of the errors associated with the individual cell samples, the standard metallurgical formulae are not sufficient. Consequently, it is necessary to utilize some type of statistical technique capable of distributing the individual cell sample errors in such a way so as to provide a physically useful mass balance for the entire circuit. This paper explores the use of linear and least squares programming as one technique to providing credible mass balance calculations. For linear programming this was done by distributing the error using two different criteria involving specific linear programming formulations: least absolute sum minimization (which assumes an underlying double exponential distribution error structure) and minimax (which minimizes the maximum absolute deviation and assumes an underlying rectangular distribution error structure). These results were then compared to the more standard formulation of minimization of the sum of least squares which assumes an underlying normal distribution error structure. The least absolute sum approach was particularly attractive due to its simplicity of use and the ease in which repetitive calculations can be done with no user intervention.

Jan 1, 1990