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By A. M. Edwards

In many ways a government could be regarded as a macro business enterprise. Its shareholders are its peoples, the productivity of the nation depends on them, and, indeed, the profits or losses of a nation's balance sheet can be linked to the surplus or deficit in its balance of payments. International loans need to be repaid, and in most developing nations the noose is tied so tightly that one wonders if such debts will ever be repaid. In their endeavour to meet such obligations, their currencies may be devalued [sometimes in frightening proportions], in the hope of boosting the export potential. However, this is a vicious circle, since exports will not be achievable unless sufficient technical expertise and infrastructure exist to address the opportunities which may have been created through devaluation. Many of the poorer countries have exportable resources of minerals, which are supplied on the international market. This may be the good news, but, unhappily, steps were seldom taken to enhance the value of the mineral products through further beneficiation and the production of high-value sophisticated products. That is where the vicious circle comes in: inadequate infrastructure is created, competent and well-trained manpower is not generated, and the result has inevitably been that the GDP slips backwards, and the nation becomes progressively impoverished. To exacerbate the problem, it is unfortunate that the more impoverished a nation becomes, the greater is its birthrate, and hence the greater the need to create wealth at the very time that the nation is becoming poorer and poorer. Within this scenario, one may well reflect that this is exactly what nations on the African Continent are experiencing, namely exploding populations, although most of the poorest nations in the world are found on this continent. Yet they often possess a superior mineral production potential. Africa still suffers from what may be called the colonial influence, in that the developed world has regarded this continent as a source of raw materials to feed its own factories. The colonial powers gave only secondary attention to developing the nation's peoples and its infrastructure, such as railways, roads, harbours, and other services. These problems all naturally fall within the ambit of government interest, and it is therefore obvious that the state should play a dominant role in addressing them. The reasons for governmental involvement in mineral-related research and development differ considerably, depending on the state of industrial development of that country. Thus, in the highly industrialised world, government's interest will lie mostly in the fields of strategic minerals, which may not be economically exploitable. An example would be the exploitation of American low-grade chromites during the Second World War, or minerals which need to be imported from unfriendly or unreliable sources, such that disruption would not be in the national interest. Colonial powers like the United Kingdom and France established R&D facilities to assist, inter alia, in the exploitation of mineral resources in their colonies. The approach of developing countries would differ from that of the industrialised nations, in that their primary quest would be to promote, support and encourage the development of infrastructure, and the establishment of major industries in order to create and provide employment opportunities, as well as to support small operators who cannot afford to undertake their own research and development. On the debit side, it is a sad fact that, however good their intentions may be, underdeveloped countries are invariably unable to provide the requisite financial support to sustain their own research-and-development laboratories. In a high-tech world, the cost of sophisticated equipment is steep and cannot be afforded. Again, the proverbial vicious circle comes into play. The competent researchers find themselves without rewarding employment and would therefore shun such institutions. Recent trends in the developed world have been to privatise research-and-development institutions, the argument being that they should draw closer to the industry they serve, and that the provision of taxpayers' money for strategic purposes is no longer as important as it may have been once. Complete state withdrawal has had its negative effect in that long-range, high-calibre research has tended to be replaced by short-term problem-solving projects. Industries engaged in the minerals business differ somewhat from major ones, like motor, computer or pharmaceutical industries, which are large enough to support their own internal research and development, and whose work is highly confidential, so that any leakage could have disastrous consequences. Furthermore, private companies are answerable to their shareholders and costly long-range, 'blue-sky' research would be difficult to justify. Since mineral research has become of lesser importance in the sophisticated First World, this trend has persisted. On the positive side, R&D institutions which are totally financed by government, suffer because they tend to forget the reason for their existence. They have often become a form of auxiliary to an academic institution where higher degrees are collected by scientific boffins who, in their quest for academic excellence, tend to disregard the benefits that industry may ultimately gain from this research endeavour. The trend in recent years - and South Africa is no exception - has been for the state to progressively reduce its subsidy, whilst at the same time encouraging research organisations [for their ongoing survival] to interact closely with industry, and to conduct work directly for them on a contract basis. Organisations such as CSIRO in Australia, CANMET in Canada, the USBM in the United States and Mintek in South Africa all generate income from the private sector. However, the critical fact is that the State should never withdraw

Jan 1, 1995

By Bernard F. Anderson

INTRODUCTION In this chapter, the term "down-the-hole drill" (DTH drill) is used as a generic name that encompasses the various trade names and other references such as "downhole drill," "in-the-hole drill," etc. This chapter is limited to a description of DTH drills used in stoping large underground ore bodies. DTH drills differ from conventional drills by virtue of the placement of the drill in the drill string. The DTH drill follows immediately behind the bit into the hole, rather than remaining on the feed as with ordinary drifters. Thus, no energy is dissipated through the steel or couplings, and the penetration rate is nearly constant, regardless of the depth of the hole. Since the drill must operate on compressed air and tolerates only small amounts of water, cuttings are flushed either by air with water-mist injection or by standard mine air with a dust collector at the collar. HISTORICAL DEVELOPMENT Mine managers have long known the economies enjoyed by quarry and open-pit operators in producing large quantities of ore. The savings are due primarily to the availability of massive equipment, capable of drilling large blastholes to reduce the amount of drilling, increase the fragmentation, reduce secondary blasting, and im¬prove the flow of the product. In an attempt to reduce underground mining costs, various methods are used for long-hole drilling, includ¬ing standard pneumatic percussion drifters and diamond drills. These systems have their shortcomings; percus¬sion drills are limited to small hole sizes and they ex¬perience excessive deviation and significant loss of energy with increased depth. The diamond drills provide deeper and straighter holes, but only at high cost. Both systems suffer from high noise levels, low penetration rates, and poor explosives distribution, among other problems. When the mining companies approached the drill manufacturers for a compact and portable large-hole jumbo for underground use, they specified not more than 1 % deviation on 60 m (200 ft) of vertical hole and a penetration rate of 15 m/h (50 fph). On Dec. 23, 1960, a test unit was placed in service in Montana and met the performance criteria. Though lacking the so¬phisticated features available today, the economies of surface blasting were brought underground. Unfortunately, the first system did not gain immedi¬ate acceptance in the industry. Among the factors con¬tributing to its demise were resistance to change, the need to alter development methods for the ore bodies, and a lack of flexibility in moving the rig from setup to setup and from level to level. In 1972, the mining industry again challenged the drill manufacturers to provide a workable jumbo that would combine compactness, ease of maintenance, relia¬bility, and efficiency, all on a self-propelled chassis. The manufacturers responded by providing improved jumbos, which have been accepted with enthusiasm throughout the mining industry. Today's DTH jumbos are capable of drilling from 100 to 200 mm (4 to 8 in.) diam holes that can be reamed to even larger diameters. The holes can be drilled to depths of 150 m (500 ft), depending upon ground conditions and the capability of the jumbo to retrieve the steel and drill. Fig. 1 illustrates a typical DTH jumbo. APPLICATIONS The uses to which DTH drill jumbos have been put are quite numerous, with new uses being found regularly. For convenience, these uses may be classified as primary blastholes and nonblasting holes. Primary Blastholes The original purpose for the development of the DTH jumbo was for drilling primary blastholes that could be mined by open-stope methods. Prior to the advent of the DTH jumbo, extensive development was required before production drilling could begin. Sub¬levels were required to allow access for column-and-arm stopers or ring/fan jumbos, to the extent necessary based on the effective penetration of the chosen machine. With the DTH jumbo, the mine engineer is able to reduce preproduction time and development costs. How¬ever, the most significant saving results from an im¬proved cost per ton of broken ore in the production phase. To utilize a DTH system, only a top heading and drawpoints are necessary. The top heading can be the width of the ore body with a 3.7-m (12-ft) back. A drop-raise pattern is drilled and shot to begin the stoping operation, providing a free face for subsequent blasting. A typical layout is illustrated in Fig. 2. The advantages of this system include: 1) Drilling and blasting are independent operations, and blasting can be performed at a rate congruous with the mine's ton-per-day capacity. 2) The development layout is simplified. 3) Good explosive distribution is achieved, provid¬ing more uniform fragmentation. 4) Environmental conditions for operators are im¬proved, including improved safety with all work directed downward (not overhead), lower noise levels, little fog, and a reduced dust count. 5) Improved production per manshift. 6) Simplified and easier operator work cycles. 7) Reduced cost per ton of product. 8) Fewer holes lost due to ground shifts. Nonblasting Holes With the introduction of the compact DTH jumbos, other practical uses became apparent, including the drilling of: 1) Holes for sand fill, from level to level and from level to stope. 2) Drain and dewatering holes. 3) Power and communications cable holes.

Jan 1, 1982

By T. N. Moore, E. M. De Souza

J. Dasher Regarding the March 1991 ME technical paper by De Souza and Moore: For more than a decade since my February 1981 article on how to use modern metric, which SME-AIME had decided to do, I have monthly pointed out metric errors to the editors. In part, I do this because there has been no action to allow editors to fix figures and tables or to allow them to require authors to do so. The latest resulting atrocity provokes this discussion of vibrating drill handle units being stated in decibels. Reply by T. Moore We have read the discussion of our paper by Mr. Dasher. Our reaction is one of surprise and incredulity. It would seem that Mr. Dasher takes exception to the use of the decibel scale to present vibration acceleration data, and the use of hertz as the unit for frequency. The basis for his objection to the decibel appears to be that it has no dimensions (which somehow invalidates its use), that it is "non-metric" and, finally, that it is parochial (of limited or narrow scope). His objection to the use of the term hertz is not stated, but we will assume that it stands condemned as "non-metric" and parochial. Obviously we disagree with Mr. Dasher's views and will now outline our reasons. Although the decibel scale originates from transmission line theory and telephone engineering, it is also at present widely used, not only in the fields of electronic engineering and acoustics, but also in the area of vibration. The original definition of the decibel (dB) was based on power ratios: dB = 10 log 10(W/W0) where Wo is a reference power. However, as the power measured across a given impedance is related to the square of the force acting upon this impedance, Z, a more commonly used definition is: [2 dB = 10 logF /Z) = 20 log F/F 10\ F0 2 /Z(0)] where F and F0 are the r.m.s. values of the forces. Now, if the measurements are related to one and the same impedance, the decibel notation in the form of 20log10(X/Xo) may be used as a convenient relative magnitude scale for a variety of quantities. Thus, X may, for instance, be an r.m.s. displacement, velocity or acceleration. It is only required that XD always be a reference quantity of the same type as X. That is, when X represents an acceleration, then X0 represents a reference acceleration. This is the formulation used in our paper. This was not an arbitrary choice on our behalf but reflects standard practice as specified in the International Standard ISO 5349-1986(E) Mechanical Vibration - Guidelines for the Measurement and the Assessment of Human Despite the metric prefix, the decibel is a parochial expression of (l) the logarithmic ratio of the loudness of a sound to what is normally audible or (2) the logarithmic ratio of two power signals in radio or electronics. A decibel is not a unit, much less an SI, unit and has nothing whatsoever to do with the acceleration of drill handles. Stating that m/s2 (acceleration) is decibels is without reason. Whoever reviewed this material should not have allowed publication of figures of dB and H.[ ] Exposure to Hand-Transmitted Vibration. This was clearly stated in the "measurement protocol" section of our paper. This quantity is then referred to as the acceleration level and is expressed in dB. We may have inadvertently caused some confusion when we simply used the term acceleration to refer to acceleration level on our diagrams. At the time, we felt the use of dB or m/s2 would make the context clear to the reader. For any confusion this decision may have engendered, we apologize. Since the decibel expresses the ratio of two like quantities, it certainly has no dimensions. It is, however, common practice to treat "decibel" as a unit as, for example, in the sentence, "The acceleration level measured at the operator's hand was 160 dB." The expression of measured quantities in dimensionless form is not inherently unacceptable. In fact, in many areas of engineering it is standard practice (consider the use of Reynolds Number, Nusselt Number, etc.). The fact that the decibel is a dimensionless quantity makes the question of whether it is a SI unit nonsensical. However, it is valid to insist that the dimensional quantities used to obtain the decibel values be expressed in SI units. A careful reading of our paper will make it clear that the measured acceleration was, in fact, expressed in units of m/s2 as was the reference acceleration (l x 10-6 m/S2). These are the accepted derived SI units for acceleration. See, for example, the standard ASTM E380-89a Standard Practice for Use of the International System of Units (SI) (The Modernized Metric System). Concerning Mr. Dasher's implication that hertz (Hz) is an unacceptable unit of measure for frequency, we would again refer him to the standard ASTM E380-89a. Here, he will find (section 2.4.2) that hertz is an accepted "special name" for the derived SI units-1. This is in keeping with numerous other international standards including ISO 5349-1986(E) to which we referred in our paper. In conclusion, we agree with Mr. Dasher on the desirability of expressing measurements in modern SI units. But we would remind him that the standards that define the use of these units, and the accepted means of presenting measured data, are in a continual state of refinement. It is, therefore, incumbent upon him to keep abreast of these changes if he wishes to constructively critique the work of others.[ ]

Jan 1, 1992

By E. Bonates

Dispatching systems have attracted considerable attention in the last years due to substantial gains in productivity achieved through their implementation. These systems will be extensively used in the future by mine operators as a means of optimizing productivity. This paper describes a computer simulation model designed to study the feasibility of continuously dispatching trucks in open-pit mines based on different operating procedures. The system developed is semi-automated and performs an active role by suggesting to the dispatcher the best assignment according to each specific circumstance. It is modeled based on an advance-clock increment to reduce computing time and allows the insertion of heuristic dispatching policies and equipment real-locations when breakdowns occur. Dispatching procedures taking into account long-term production objectives are considered in system development.

Jan 1, 1988

By Keith A. Yarborough

INTRODUCTION The National Park Service (NPS) conducted a research program from mid-1975 to early 1978 to measure alpha radiation levels in natural caves which it administers. Subsequently, a long-term management program was developed which has conducted routine monitoring of radiation levels and has computed employee exposure accumulations in certain NPS caves. The overall program: research and management; was undertaken to evaluate the possible health hazard to cave visitors, interpreters, researchers, maintenance personnel, concessions employees and other workers and to protect their health. The results of this work have been reported extensively elsewhere (see References 1 through 3, 5, 6, 9 through 13, and 14 through 26). This paper deals with the relation of radon gas and daughter produced radiation levels to the cave air flows which mobilize them. These flows are a function of interior cave and exterior ambient air temperatures and pressures and of the cave's geophysical configuration. The low-level, ionizing radiation is produced by minute concentrations of radon and thoron gases which emanate from within caves. Because of confinement, the radiation levels are often appreciably higher than in surface atmospheres. Subsequent radioactive disintegration of the gases into their ionized "daughter" products, which are particulates, permits distribution of the alpha radiation throughout a cave system. The extent and character of this distribution depends upon the particular cave and the air flows which it produces. Thus, the alpha radiation serves as a "tracer" to describe the flows. The flow analysis is divided into two parts: 1) A qualitative description with respect to the two basic geophysical cave types over both long-term (annual) and short-term (diurnal to weekly) periods. 2) A quantitative description of the unsteady, uniform and non-uniform, one-dimensional, incompressible motions for both influent and effluent air flow situations in both basic geophysical cave types. A part of the qualitative description presents "Special Measurement" data: Tsivoglou [vs.] Kusnetz working levels, Tsivoglou individual daughter levels and free ion concentrations, radon gas concentrations, and equilibrium ratios. A great deal of important work has been carried out by Mr. Bobby C. Carson, Cave Radiation Technician at Mammoth Cave National Park, Kentucky. He reports these findings later in this conference (Ref. 6). Some of his results have been used here to establish the cave air flow analysis. Other National Park Service personnel have made measurements at others of the caves for which data are reported here. All of this work and cooperation has been vital to the success of this research program. It is very much appreciated. THEORETICAL FLOW DESCRIPTION Previously reported research (See Ref's. 20, 21, 23 and 25) has established the qualitative relationship between the alpha radiation in caves and their natural air flows. The radiation serves as a representation of these air flows. Changes in radiation with time represent changes in the main forces which produce the air flows. The quantitative data have substantiated that [all caves in which the primary cause of air flow is due to temperature produced gravity (density) gradients and also having minimal man-made disturbances, experience seasonal variations in airborne alpha radiation. The radiation levels increase in summer but decrease in winter], based upon seasonal air movements through each cave system which occur naturally. Two general types of physical cave configurations which control the air flows have been identified as: 1. Those which go up into a hillside or mountainside: Type I ("Upside-down" = USD). 2. Those which go down into the earth; Type II ("Right-side-up" = RSU). These act to control the air flows seasonally. The summer increase in Type I caves is due to increased air flows, whereas in Type II caves it results from stagnation or reduced air flows. This seeming paradox is explained by the physics of the air flow regime in each cave type and has been detailed elsewhere [Ref's. (20), (22) through (25)]. It[ i] true, in general, that air flow decreases airborne radiation in the [immediate vicinity] in which it occurs. Any paradox results from subsequent distribution throughout the case as to [ how] the air moves through a cave [system] with respect to time and space. Exceptions to the cave air flow "rule" are: 1. Caves in which pressure gradients and pressure fluctuations [predominate] in producing the air flows. 2. Caves in which man-made effects and management practices are superimposed on the natural air flow regime. Man-made disturbances which can alter the natural cave air flows are tunnels, elevator shafts, bore holes, sealed and closed portals, etc. If these are not properly sealed, the natural air flows which they change will totally alter the distribution and seasonal variations of the alpha radiation levels. These exceptions may act separately or in combination.

Jan 1, 1981

By William Lyman

GENERAL DESCRIPTION Shrinkage or shrinkage stoping refers to any mining method in which broken ore is temporarily retained in the stope to provide a working platform and/or to offer temporary support to the stope walls during active mining. Since ore "swells" when broken, it is necessary to shrink the muck pile a corresponding amount by draw¬ing some of the broken ore out as the stope is advanced-hence the name. Broken ore retained during stoping is drawn out after the stope has reached its limits. The stope may be left empty or may be filled with waste contemporaneous with, or subsequent to, the final draw. Traditionally the method implies conventional overhand stoping methods with miners working between the muck pile and the stope back, in a space which advances updip with mining and is maintained by balanc¬ing "swell" with "shrink." The shrinkage classification is also applicable to so-called "semishrinkage" methods in open pillar-supported stopes where broken ore is temporarily retained as a working platform but offers no wall support; and to various blasthole shrinkage methods which utilize broken ore temporarily retained in the stope for wall support, but which do not require miners to work from muck pile in the stope. The method is generally applied to steeply dipping veins of strong ore between strong walls. APPLICATION Geometry The geometry of a shrinkable vein is described in terms of dip, width, and regularity along dip. Overall strike and dip dimensions and irregularities along the strike generally impose no restrictions on the method. Dip is ideally 1.2 to 1.5 rad (70 to 90°). As dip falls below 1.2 rad (70°), the shrinkage draw begins to strongly favor the hanging wall side, thus leaving a poor working platform for conventional overhand work. This is particularly true in relatively wide stopes. The sup¬port afforded to the hanging wall also diminishes with decreasing dip, reaching nil as the dip approaches the repose angle of broken ore. Dips below 0.78 to 0.87 rad (45 to 50°) are not generally shrinkable except by open stope "seinishrinkage" methods. Minimum mining width is fixed by working space requirements in the stope-generally about 1 m. Shrink¬age in narrower veins requires that waste rock from one or both walls be broken with the ore and the attendant dilution accepted to achieve the minimum width. Nar¬row stopes are less suitable, encouraging hang-ups and bridging of broken ore, with the attendant problems of erratic draw and incomplete recovery of broken ore. Maximum practical width may be 3 m or less to over 30 m, depending upon the competency of the ore and its ability to stand unsupported across the stope back. This is a vital safety consideration in conventional over¬hand stopes, but is much less of a factor in blasthole shrinkage methods. Very wide veins and massive ore bodies have been mined by transverse vertical shrinkage panels separated by transverse vertical pillars which are either abandoned or recovered later by other methods. Regularity along the dip is a prerequisite of shrink¬age as there must be no serious obstruction to the flow of broken ore downward through the stope to the sill level. Gentle rolls along the dip are acceptable if the local footwall dip everywhere exceeds 0.78 to 0.87 rad (45 to 50°). Off-dip hanging wall and/or footwall splits can generally be mined selectively from a conventional shrink stope as they are encountered without ad¬versely affecting subsequent continuation of shrinkage mining updip on the main vein. Vertical offsets or major rolls along the dip which cannot be "smoothed over" generally require that a sublevel be established with new draw control development. Blasthole shrinkage methods are much less flexible (and thus less selective) in their ability to accommodate any of these irregularities. Ground Conditions The wall rock must be strong enough to stand with the minimal support afforded by the dynamic mass of broken ore in the stope. During active mining, local sloughing from the walls is restrained, but the broken ore affords little, if any, useful resistance to closure of the stope walls. Such squeezing, if present, may bind up the stope and cause the loss of much ore. Pillars left between and/or within stopes are effective in preventing closure but reduce overall recovery. Walls may be re¬inforced by bolting after each stope cut in conventional shrinkage but not in blasthole shrinkage. Ore in place must be strong enough to stand with no natural support across the stope width, although tem¬porary artificial support or reinforcement may be used locally in conventional stopes. Some spalling or sloughing is permissible in blasthole shrinkage as men are never present in the stope. Physical and/or mineralogical characteristics of the broken ore may impose restrictions on stope design and/or operational plan¬ning, and may even preclude the use of shrinkage al¬together. Examples include: ores which, when broken, are cohesive or which tend to pack or cement together under the influence of ground water, wall pressure, and/ or chemical reaction. Such conditions precipitate er¬ratic draw during mining and often result in difficult and/or incomplete final draw; pyritic ores which oxidize very rapidly in the stopes and may generate heat, imposing a fire hazard by spontaneous combustion; sulfide ores which oxidize sufficiently in the stopes to adversely affect mill recovery by flotation; and ores (es¬pecially those containing uranium minerals) which ex¬ude radon gas and thereby impose ventilation constraints on stope design. In most cases these problems can be minimized by limiting the size of stopes, by minimizing the duration of mining activity in each stope, and by promptly drawing each stope empty following comple¬tion of mining.

Jan 1, 1982

By R. Vogt, H. C. Mult, F. L. Wilke

INTRODUCTION At present, the capability of production planning software based on Linear Programming (LP) is still limited to the optimization of the single LP-run. This is due to the LP-model itself which cannot consider the interdependencies between individual LP- runs. With regard to planning of selective mining this limited way of optimization frequently leads to situations, where the remaining and accessible ore blocks do no longer allow to produce ROM-ore in the qualitative composition required by the ore processing plant. However, many of the aspects taken into consideration when setting up production plans built from mutually dependent LP-runs cannot be modelled in a system of linear equations. They are thus unsuited for treatment with LP and have to be taken care of by the planning engineer without any assistance by the system. The KBS currently under development is intended to assist the planning engineer in designing a production plan under special consideration of the combination of consecutive LP-runs and blending beds. It is not necessarily intended to find the optimum solution within a given planning situation which is, anyway, hard to determine due to the multitude of influences. The objective is rather to work out a good and - from the practical point of view - feasible production plan. The new aspect with respect to mine planning is the integration of expert knowledge and experiences via the KBS into the planning process in order to support the planning engineer. The planning system is being developed in close cooperation with an iron-ore open pit mine. COMPONENTS OF THE PLANNING SYSTEM The software is being developed on a workstation under UNIX and comprises the components LP, CAD-module and the KBS. The applied multi-goal LP-algorithm is an in-house development of the Department of Mining Engineering at Technical University Berlin. It was already successfully implemented within other mine planning programmes and was only slightly adapted to the specific needs of the present system. Within individual LP- runs it finds the optimum qualitative composition of ore production in the sense of the selected optimizing criterion and under the given restrictions: i.e. it determines tonnages to be mined from blocks in order to optimally meet the requirements of the ore pro- cessing plant. A CAD-module based on the commercial SURPAC package in combination with a simulation device is used to graphically depict the block model and progress of mining. Both LP-algorithm and CAD-package are integrated in the KBS. It has been decided to use the shell NEXPERT OBJECT as it is a hybrid system which supports both rule-based and object-oriented knowledge representation. MINE-MODEL AND LP-MODEL KBS have to be tailor-made for specific planning problems. Therefore, it had to be decided which specifications of the iron-ore mine should be represented in the model. As the limited possibilities of a university institute do not allow to develop a KBS for mine planning which is ready to use in industry, it was decided to concentrate on those characteristics that can be regarded as typical for iron-ore surface mines and that seemed to be suited for treatment with knowledge-based techniques. The following chapter summarizes the most important features of the mine model. The description of the requirements to the mine's sales products is followed by an outline of the applied LP-model. Mine model • The model of the mine as it is used for planning consists of • the block model of the deposit, • the mobile equipment, • stockpiles and blending bed and • the requirements to the sales products. The deposit is described by a block model which contains data on the chemical composition, LOI, grain size and tonnages. Grain size was included as it is important for the two sales products of the mine. Furthermore, it is known whichs blocks require and which don't require blasting; this is relevant to the assignment of loading equipment to individual blocks. The blocks are devided in three categories: • ore, which will directly be taken to the blending bed; • waste, which will be put on the waste dump; and • blocks which will be either transported to the blending bed, to stockpiles or to the waste dump depending on the specific planning situation. This decision is made during planning. Neighboring blocks are combined in mining areas to which the loading equipment is individually assigned. Mobile equipment comprises shovels and wheel-loaders as well as trucks. The characteristics of the loading equipment are important for their ability to load different blocks and for the permissible degree of their re-positioning etc. The mine disposes of a blending bed for homogenization of the production, of a waste dump, and of several stockpiles with different ore qualities. The requirement to make only limited use of the stockpiles for economic reasons is included in the KBS. According to long term planning two commercial products have to be produced, which differ both in grain size and qualitative composition (TABLE 1). Their mass-proportions in the blending bed have to be within a fixed range. Not considered in long term planning is the occasional need for lump ore, which occurs at very short notice and has to be produced in a "campaign-like" manner. This requires the total re-arrangement of all plans for on- coming blending beds and would therefore be ideally suited for

Jan 1, 1996

By E. Fahrenkrog

An intelligent supervisor developed using expert system tools and applied to a crushing plant is presented in this paper. The structure of this supervisor, based on a model of a human operator, consists of a Knowledge base containing a qualitative model of the process, and a Rules base with the heuristic logic used by an operator when supervising a plant. The simulation results obtained show the potential of this type of systems to supervise the automatic operation of a plant.

Jan 1, 1989

By G. D. Bates

INTRODUCTION The Magmont mine is a joint venture of Cominco American Inc. (operator) and Dresser Minerals, Inc. The mine-mill operation is located approximately 160 km (100 miles) southwest of St. Louis, MO, in what is commonly referred to as the "Viburnum Trend.” The Magmont mine is designed for a production rate of 3810.2 t/d (4200 stpd) on a 5-day week, three shifts per day basis. Initial production began in 1968. The mine is open stope, room-and-pillar, and essentially horizontal along the trend of the ore body. Briefly, the main geological structure can be described as a brecciated graben bounded by reverse faults. The ore body in cross section is shaped like a bell curve with some lateral extension at the lower part. Presently outlined ore is 609.6 to 762 m (2000 to 2500 ft) in width and 2133.6 m (7000 ft) in length. The ore varies in thickness from 4.87 m (16 ft) on the fringes to an average of 27 m (90 ft) in the high ore areas bounded by the reverse faults. Lead is the primary metal with zinc and copper secondary. MINE DESIGN The basic design of open stope, room-and-pillar mines has been described by several writers and need not be repeated here. (Anon., 1970; Bullock, 1973; Casteel, 1972; Christiansen et a]., 1970; and Lane, 1964) This discussion covers the mining sequence as applied to the particular problems at the Magmont mine, the use of equipment, and deployment of the work force. In the upper portion of the Magmont ore body is a layer locally called the False Davis shale. This layer lies below the true Davis shale, is normally interbedded with dolomite, is of varying thickness, and if mineralized, is included in the top pass of the mining sequence. In thick ore areas this layer will be 2.13 to 2.43 m (7 to 8 ft) in thickness and will occur in the upper portion of the pillars. Due to its incompetency the presence of this False Davis layer is of primary concern in mine planning and operation. Mining areas are divided into three basic groups by ore thickness. First are areas of ore up to 6.09 m (20 ft) in thickness. These areas are below the False Davis shale and are mined single pass with drill jumbo. Second are those areas up 13.71 to 15.24 m (45 to 50 ft) in height. The first 4.87-111 (16-ft) Pass is taken at the top of the ore and the back and pillars secured. Benching the lower portion(s) in 4.57 to 4.87-m (15 to 16-ft) passes is then done with either a drill jumbo drilling horizontally or a crawler drill drilling vertically. Normally these areas are below the Table 1. Productivities per Manshift False Davis shale. These areas may also be mined by back slashing, or overhand benching, where the first 4.87-m (16-ft) pass is taken at the base of the ore and successive 4.87- m (16-ft) passes are taken upward. A minimum amount of back slashing is done at Magmont since it requires repetition of roof control on each pass and roof control is the single largest stoping cost at Magmont. Ore left to provide a working platform oxidizes and is coated by oil spills thus reducing metallurgical recoveries. The third mining area is over 15.24 m (50 ft) in height UP to a maximum of 40.23 m (132 ft) and will encompass the False Davis shale. These areas are mined by first driving +15% inclines to the top of the ore body. The top pass is mined and the back is bolted and roof mats installed as a matter of standard practice to minimize roof problems as mining progresses downward. Once the back and pillars on the top pass are secured, benching begins on successive passes with either the drill jumbo or crawler drill. Pillars on all successive passes below the top pass are secured as necessary. While benching progresses below the top pass, the pass at the base of the ore body is mined leaving a sill of 4.57 to 7.62 m (15 to 25 ft) in thickness to be removed with the crawler drill in a retreating manner. Rooms are mined on a 1.57 rad (90") grid pattern to insure alignment of pillars where multiple passes are taken. Pillars are designed on a 17.98-m (59-ft) spacing with rooms up to 10.66 m (35 ft) in width. Heading widths are wide enough for the mobile equipment to turn without additional allowance for curves. The result is a flexible layout which provides a maximum number of headings available for high extraction rates and grade control. PRODUCTION Incentive Bonus Incentive bonuses play an important part in the mine production at Magmont. Production crews are trained to perform only one of the mining functions of drilling, blasting, mucking. or roof bolting. This specialization, or functionalization, is augmented by development to open all possible stoping areas as early as possible in the life of the mine. This insures that each crew will have enough headings to perform its specialty. The incentive bonuses increase exponentially as output increases. The lucrative incentive bonus coupled with the specialization of the production crews and proper mine development have combined to give the high productivities shown in Table 1. Development crews perform all mining functions in their area. The incentive bonus is paid on a per foot basis, Crews on different shifts working the same heading share equally in the bonus proportional to their contract hours worked.

Jan 1, 1982

By D. T. McMurray

INTRODUCTION Shabanie mine, situated some 180 km east of Bula¬wayo, has been a producer of chrysotile asbestos for more than 50 years. The ore bodies occur in serpentinized dunite, which overlies talc-carbonate schist. A zone of relatively competent rock of varying thickness occurs between the schist and the ore bodies, which are gen¬erally less competent. The hanging wall of the ore bodies is economic, and the hanging-wall serpentine carries a variable subeconomic amount of fiber. It is important to note that, in general, the ore body competence is less than that of the foot and hanging-wall formations. Historical After surface operations ceased, cut-and-fill stoping was used to win ore from underground; this was success¬ful until the increasingly stoped-out area caused insta¬bility in the stope pillars and back. Consequently, dur¬ing the early 1950s, a gradual change to cave-mining methods was made, the ore being won by hand lashing in drawpoints, situated in the basement of the stope blocks, and passed through orepasses under gravity to the haulage level some 13 m below. About this time, interest was focused on the sub¬level caving method in use in Swedish iron ore mines: it was felt that it might be applied economically to the Shabanie ore bodies. Accordingly, in 1958, an experi¬mental stope block was laid out in which sublevel inter¬vals and extraction tunnel spacing were 9 m. The tun¬nels (ring drilling drives) were oriented on strike-in contrast to the Swedish system, in which crosscuts that retreat from hanging to footwall are used. The advantages of the method were quickly appre¬ciated by the operating personnel and, despite the in¬evitable teething troubles pertaining to the introduction of any new mining method, it was not long before sub¬level caving was providing a high proportion of the mill feed. The disadvantages also became apparent at an early stage, however, and, from that time to the present, continuing modifications have been made to mining lay¬outs in an effort to improve ore recovery. GENERAL DESCRIPTION OF METHOD The mine is served by a vertical hoisting shaft, in which two skips, a man cage and a service cage, provide adequate capacity for production requirements. The rock hoist is a Ward Leonard control hoist, in which two electric motors drive a common gearbox. The man winder is driven by an a-c motor. Several auxiliary shafts provide secondary egress and intake and return ventilation. Main haulage levels are above (Fig. I a and b). Blasthole fan patterns are drilled by drifters of 100 mm bore, drilling 41-mm holes; when a sufficient strike length has been drilled, a slot is cut in the upper¬most sublevel and the rings are broken into the slot. Initially, a limited tonnage is drawn, since it is essential to ensure that the hanging wall caves behind the retreat¬ing stope face. Once this has been established, maxi¬mum tonnage can be drawn, as described later in this chapter, under the heading "Draw Control." The broken rock is loaded by 0.14 and 0.20-m3 load¬ers into cocopans (rocker-dumping type of tipping truck), which are hand trammed to orepasses, discharg¬ing on the haulage where 11-t electric trolley locomo¬tives haul 3.95-m3 Granby cars to the main shaft bins. As is evident from Fig. 1, the layout is simple, the block is brought rapidly into production, there is a high degree of selectivity and flexibility, and the result is a low-cost high-productivity mining method. DEVELOPMENT Main haulages are developed at 3.2 x 3.2 m, and once the service winze connections have been completed the development of the sublevels is undertaken. The footwall drives are cut first, to obtain access to the block. These ends are of the standard section, 2.4 x 2.8 m, and from them crosscuts at intervals of 70 m are driven through the ore body to the hanging wall. These crosscuts are used to supplement the geo¬logical information previously obtained from diamond core drilling, and they provide additional and more de¬tailed data on fiber percentages and lengths, structural features, and other relevant criteria which are used to build up the geological assessment of the area and to classify it in terms of the geomechanics rock classification (Laubscher and Taylor, 1977). The crosscuts also allow the necessary orepasses to be sited conveniently so that tramming distances from the loading points are not excessive. Development Drilling Once the skeleton development has been completed, the extraction headings are developed at 2.4 x 2.4 m as shown in Fig. 1. Standard development practice is to use crews of a machine operator and his helper, equipped with air-leg mounted jackhammers, to drill rounds of 1.8 m with integral tungsten carbide tipped drill steel. The round drilled is a normal drag round, as shown in Fig. 2, but considerable attention is paid to the drilling of the perimeter holes to use effectively the

Jan 1, 1982

By William H. Wiese

Anecdotal reports of high rates of congenital malformations and spontaneous abortions at the Shiprock Indian Health Service Hospital in San Juan County, New Mexico prompted an interview survey, obtaining data from families of 26 former uranium miners and 30 controls. Results supported the possibility of increased fetal wastage and congenital malformations. However, definitive conclusions were obviated by methodologic limitations in terms of selection of cases and controls and interviewing technique, as well as very low numbers and lack of statistical significance. The Council of Environmental Quality (1980) published a report which singled out San Juan County, nationally, as having an extra-ordinary high rate of congenital anomalies reported from birth certificates in the years 1973-75. San Juan County is in the northwest corner of New Mexico and includes the northeastern edge of the Navajo Indian Reservation. It has been the site of uranium mining and milling and also of coal mining (Clement Associates, 1980). While much attention has been focused on lung cancer and chronic lung disease as health hazards to uranium miners, virtually no attention has been focused upon possible reproductive effects that may result from exposure to radiation. Could these reports indicate such an additional arena of health effect either on miners or on others living in the vicinity of the mines and mills? The biologic and epidemiologic literature does not lend much support for adverse genetic outcome. The UNSCEAR report (UNSCEAR, 1977) summarizes much of the existing data. For example, on the basis of many animal studies and what has been observed in atom bomb survivers in Japan, it has been estimated that mutation rates resulting from low LET radiation to paternal germ cells would be 60 X 10-6 recessive and 20 X 10-6 dominant mutations per gamete per rad. There would be a five-fold increase in the number of recognizable abortions. The doubling dose for genetic disorders in mice is estimated at 100 rads. The assumption has been that low levels of exposure, such as might be received occupationally from uranium mines and mills or environmentally will produce reproductive effects that would defy distinction from background rates. However, the possibility can not be dismissed. In man, spermatogonia are thought to be three-fold more sensitive than in the mouse. Also, the high LET radiation from alpha particles is as much as 20-fold more effective in inducing changes in meiotic and post meiotic stages of cells, compared with low LET radiation. There is some degree of localization of alpha emitters in testicular tissue after inhalation of radon and radon daughters (Blanchard, R.L. and Moore, J.B., 1971; Pohl, E., 1964). The slopes of doseeffect curves are affected by type radiation and rate of administration. Data on such curves are poorly defined and debated as to whether there may be increased or decreased effect per unit dose at low levels of high LET radiation. Increased rates of aberrations of chromosomes in peripheral lymphocytes cultured from uranium miners have been reported (Brandom, W.F., et al., 1978). At cumulative exposures of 100 working level months (WLM) or less, a range of exposure lower than the limit of current occupational standards, the aberrations were increased more than two-fold. Our interest has led us to pursue the question of reproductive effects along three separate paths of inquiry: studies of the secondary sex ratio, cytogenetic study of human sperm, and studies of rates of congenital anomalies. I will present the thrust of our findings, most of which remain preliminary, comment on their interpretation and limitations, and indicate directions needed for future research. The secondary sex ratio is the ratio of males to females at birth. In the U.S., the ratio averages around 1.05. Ratios in blacks are somewhat lower (1.025) than in whites (1.053). For American Indians, the data are less definite. Values from Oklahoma and California are similar to whites (1.05). Southwestern tribes have, as we shall see, varying levels. The theoretical effects on sex ratio resulting from irradiation of the zygote are based on the principles of sex-linked genetics, including differential effects of lethal dominant and recessive mutations on the sex chromosomes, complicated by possible non-disjunctional events. Observations have done little to clarify, confirm or refute the theories. The sex ratio of progeny of survivers of the atomic bomb showed little change (Schull, W.J., Otake, M., and Neel, J.V., 1981). Observations in children of uranium miners having been inconsistent. Most frequently cited is a study of miners in Czechoslovakia (Müller, C., Razicka, L., and Bakstein, J., 1967). Compared with children born to miners prior to the start of mining, the sex ratios of children born after the start of mining were decreased -statistically significant after correcting for birth order. The ratios in first-born children were, before the start of mining, 1.118, and, after the start of mining, .713. In the interview survey at Shiprock, the sex ratio in families of non-miners was 1.03, and in families of miners it was 0.73 (p < .05). Dr. Alan Goodman at the University of New Mexico has studied trends in the sex ratio in the general population in New Mexico and Arizona with particular reference to counties in the Four Corners area and the Navajo Indian Reservation, areas where uranium mining has been active_ His findings (unpublished) are as follows: (1) Beginning in the early 1950&apos;s, there has been a modest, but persistant and statistically significant decline in the secondary sex ratio for the state of New Mexico compared with the U.S., a decline that roughly parallels the emergence of uranium mining in the Grants mineral belt in northwestern New Mexico. (2) Data available for the 11 years since 1969 show that among the five counties with ratios significantly lower than the state

Jan 1, 1981

By S. Odell

The Bureau of Land Management (BLM) has constructed two biochemical reactor systems (BCRs) to treat mining influenced water from the Forest Queen Mine near Silverton, CO. The bioreactors coupled with a natural wetland provide treatment to the Forest Queen Mine drainage. The central objective of this paper is to review the performance of the bioreactors and wetland treatment system. The bioreactors have been in operation for nine years and along with the natural wetland are still removing a high percentage of the influent metals.

Jan 1, 2010

By Robert H. Merrill

INTRODUCTION In most cases, the design of an underground mine is based upon the premise that the ground either will cave or will be stable. This chapter concerns the design of a mine in ground that will cave readily or with some as¬sistance, such as by long-hole drilling and blasting. Some of the more widely used caving systems of mining are panel caving, block caving, sublevel caving, and large pillar recovery. Some of the less widely used systems are glory-hole, top slicing, and induction caving. Al¬though the common practice of pillar robbing is not usually considered to be a caving system, this subject will be treated as a part of this chapter. BASICS OF CAVING Caving systems are most successful in ground that will cave in sizes that will flow through openings and grizzlies, and will easily load in cars or on belts for haul¬age. The ground most likely to cave well is highly frac¬tured and contains breaks, flaws, or other discontinui¬ties that form planes of weakness. Also, caving action can be greatly enhanced if the host rock itself is low in compressive, shear, and tensile strength. Ideally, a cav¬ing system of mining is best employed when the criteria for caving is a feature of the ore body and the develop¬ment drifts, haulageways, and drawpoints can be mined in a highly competent rock beneath the mineralized zone. However, the development is often in the same, or similar, fractured rock and the openings require sub¬stantial artificial support to assure stability. Several clues can be assembled to identify potential caving ground; however, for borderline cases, no sure method has been devised to date. The diamond-drill cores taken for exploration can provide an excellent clue provided drilling is performed carefully by experienced drillers. For example, if the ground is cored in such a manner that the breaks in the core are caused more by failure of the rock than by whipping core barrels, plugged drill bits, or other drilling causes, and the intact core lengths are consistently long [say, 0.6 to 3 m (2 to 10 ft) of unbroken core], there is little reason to believe the ground will cave without considerable as¬sistance. This is especially true for rocks with compres¬sive strengths above 34.5 MPa (5000 psi) and tensile strengths above 2.1 MPa (300 psi). On the other hand, if core recovery is low (below 80%) and the recovered ore is broken in small pieces and the breaks are along obvious weaknesses in the rock, the chances are excel¬lent that the ground will cave. This is true even when the rock between the defects has high compressive and tensile strength. Another clue has already been mentioned, that is, the measurement of the physical properties of the rock and the natural planes of weakness or defects in the rock. The planes of weakness in the rock can often be detected from outcrops, cores, or other exposures of the rock under consideration. Some rock types are known to be strong and will sustain large, unsupported open¬ings and would be difficult to cave intentionally. Yet the same rock type can also contain unbonded or weak planes of weakness or fractures, and in these locations the rock would undoubtedly cave with little assistance. Therefore, although the inherent strength of the rock is a factor in caving, the natural defects in the rock are more often the deciding factor. DESIGN CONCEPTS For the most part, the design of openings for caving ground is a problem of the interaction of openings over a relatively large area of the mine. To illustrate, Fig. 1 is a simplified section of a series of openings along the grizzly level or draw level of a block caving or panel caving development, and above this opening is a simpli¬fied section of a room-and-pillar arrangement on the undercut level. At this stage of the development, the stresses around the openings on the grizzly level are only moderately influenced by the openings on the undercut level and vice versa. Therefore, the stresses around the openings are approximated by the stresses around single or multiple openings in rock, the values of which are de¬scribed in the literature (Obert, Duvall, and Merrill, 1960; Obert and Duvall, 1967). Once the pillars on the undercut level are blasted (Fig. 2), the situation changes abruptly. The undercut opening (prior to caving) now can be approximated as an ovaloidal opening above the grizzly drifts and this opening tends to shield the vertical stress field. As the caved stage is drawn the stope approximates a much larger rectangular or square opening filled with rock, and if the rock is not sustaining a major portion of the stress field, this opening can be considered (for en¬gineering purposes) to be empty and the stresses that interact between the larger and the smaller openings take on a totally new perspective (see Fig. 3). Next, let the material cave to the surface, and let the caving ma¬terial sustain some stress, but much less than if the ma¬terial were intact. This condition is similar to a soft inclusion in a rigid body and has been treated in the literature (for example, Donnell, 1941). At this point in time, the grizzly drifts are subjected to the stress con-

Jan 1, 1982

By S. J. Schatzel, K. M. Ajayi, E. N. Rubinstein, Z. Khademian

This study simulates the impact of a shale gas well casing breach near a longwall mine. Field studies are conducted to measure mining-induced permeability changes over the abutment pillar of a longwall mine, and a geomechanical model is developed in 3DEC, a three-dimensional numerical modeling code, to predict the aperture of fractures in the overburden at the study site. The predicted aperture values are used to determine mining-induced permeabilities, and the results are compared with the field measurements. These aperture values are provided as inputs into fracture flow code (FFC), which generates a stochastic discrete fracture network (DFN) model for the study site and predicts the potential shale gas flow to the mine. Results from 100 DFN realizations are statistically analyzed using the bootstrapping method to compensate for notable variation in fracture geometry. The results show a significant difference between the gas inflow for nearby panels due to increase in the induced permeability during mining of the second panel. The average gas flow to the mine was calculated as 4.72 × 10 m3/s (49 cfm) for a hypothetical breach at the Sewickley horizon during the first panel mining, 8.97 × 10 m3/s (19 cfm) for a hypothetical breach at the Uniontown horizon during the first panel mining, 2.16 × 10 m3/s (458 cfm) for a hypothetical breach at the Sewickley horizon during the second panel mining, and 8.07 × 10 m3/s (171 cfm) for a hypothetical breach at the Uniontown horizon during the second panel mining. Depending on the mine ventilation system, this could result in methane concentrations exceeding regulatory limits. Hence, these findings provide insights into the potential risk of an unconventional gas well casing breach near a longwall mine under shallow cover.

Feb 1, 2024

By Morris Beattie

Arseno Processing Ltd. has developed process technology for the extraction of gold from refractory ores and concentrates. These ores, which do not respond to conventional extraction processes, are rapidly becoming the primary targets for mining companies wishing to become significant gold producers. It has recently been estimated that by the mid 1990's, the annual gold production from the Pacific Basin Epithermal Arc in southeast Asia could be as high as 550 tonnes. Within this area the Porgera deposit under development by Placer, Renison and MIM, and the Lihir Island deposit under development by Kennecott have been shown to be refractory and require pretreatment before the gold can be extracted. The additional deposits which are now starting to be developed in this area are also expected to require such pretreatment. The Arseno process promises to be an economical pretreatment for these ores and to result in the maximum achievable gold recovery. Numerous refractory deposits are also known in Australia, North America, China and other parts of the world. Samples from many deposits throughout the world have been tested using the Arseno process conditions and the consistently successful results have demonstrated its widespread application. The samples which have been tested include ores in which the gold is locked in pyrite, marcasite or arsenopyrite. The process is insensitive to the sulphur content of the feed. Feed materials containing from 1 % to 50% sulphur have been tested successfully with the process. In the Arseno refractory gold process, sulphide minerals are dissolved so that the contained gold becomes accessible for subs7qu7nt gold recovery. This gold recovery is achieved by cyamdat1on of the process residue. For many of the materials which have been tested, gold extractions in excess of 95% have been achieved.

Jan 1, 1998

By S. F. Walsh

The San Manuel Oxide Pit is located within the subsidence boundaries of Magma Copper Company&apos;s block caving operation. Ground conditions offer a constant blasting challenge and vary from hard, competent crystalline rock to widely spaced fracturing in a sedimentary conglomerate. A successful drilling and blasting program has undergone several changes since mining commenced in 1985 and continues to be developed. This paper discusses the site specific progress made with the drilling and blasting program from selection of feasibility parameters to current practice. Topics discussed include controlled blasting, use Of high energy blasting agents, charge decking, blast casting, as well as standard drill and blast techniques.

Jan 1, 1988

By B. L. Seegmiller

Regular entry trusses are used in headgates, tailgates, roadways and intensely-faulted areas. These trusses have proven to give the required support and maintain roof stability even under major abutment loading conditions for both the re-treating and advancing longwalls. Areas previously requiring steel arches have been eliminated. A special Continuous Entry Truss? provides permanent support to adjacent tunnel/crosscut zone for more than 300 feet in each tunnel. The continuous nature of the tensioned truss allows sup-port forces to be placed on the roof in not just the two-dimensions of regular trusses, but in a full three-dimensions. A newly developed Roktruss? provides support in an arched 30-foot high rock zone. The Roktruss? allows truss support concepts to be used in an opening with a curved roof.

Jan 1, 1989

By Ryan Dr.

In December 1988 Phelps dodge corporation brought an underground gold mine into production in the republic of Botswana. The mineralization is associated with steeply dipping shear zones in Archean greenstones of the Tati Schist Belt. The gold is intimately associated with sulfides and is refractory in nature. The deposit is being exploited by underhand open stoping methods. Vertical and inclined shafts are used to hoist the ore to surface where it is treated by conventional crushing, milling and flotation, followed by fluid bed roasting with CIP elution and carbon regeneration. Adjacent deposits are being explored with a view to increasing production. Legal and taxation aspects of operating in Botswana are discussed.

Jan 1, 1989

By J. W. Leonard

Breakage, handling and treatment techniques that result in the placement of greater quantities of coal in fixed volume haulage units such as ships, railroad cars, trucks, mine cars, shuttle cars and scoops can do much to increase productivity by improving bulk density to reduce delivered cost, surprisingly enough, only in steel industry by-product coking has this subject been investigated and applied to practice, where as much as 20% or more productivity has been achieved. Research at UK Mining using popular coal mine loading cycle simulators shows that for typical Central Appalachian mining conditions a 10 to 20 percent gain in bulk density can produce a 15 to 30 percent increase in the loading cycle time of fixed volume shuttle-car/scoop based mining systems.

Jan 1, 1989

By C. Williard

Computer based mine monitoring and control systems represent a relatively new technology to be applied in the U.S. mining industry. While computer assisted engineering, design, and planning have been accepted as tools for specific technical and engineering jobs within the industry, integrating mine-wide monitoring and control systems into mining operations has been slow. This paper reviews the state of the art of mine monitoring and control systems and provides insights into future trends and developments.

Jan 1, 1988