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By W. C. Yates

Lack of visibility from cabs of the large haulage trucks used in open-pit mines has become a serious safety problem. This Anaconda research program was conducted to develop a low-cost system for alerting haulage truck operators of the presence of small vehicles in their blind areas. Several systems, developed and tested to help solve this problem, are summarized and evaluated in this paper. As the size of the haulage trucks used in large open-pit mines has increased, so have the blind areas. Mine utility vehicles, pickup tucks and cars become concealed within these blind areas and, thus, become a serious safety hazard. Figure 1 shows the typical blind areas of a 150-ton haulage truck. The areas of greatest concern are those at the rear center and front right corner of the truck. It is within these areas of "no visibility" that numerous small vehicles over the years have been crushed by the large haulage trucks at various mine sites. Although the occupants in most instances, have escaped the vehicle before it was destroyed, injuries and deaths have occurred. The accidents, in most cases, were attributed to negligence or carelessness on the part of the driver of the smaller vehicle, but the problem of limited visibility still exists, and a solution is needed.

Jan 1, 1982

By James W. Baxter

State geological surveys promote sound procedures for evaluation of mineral resources before development-procedures that protect the environment and accrue maximum benefit to citizens. Sophisticated laboratories have been equipped for geophysical, geochemical, petrographic, and mineralogic characterization of materials to assist citizens in land-use planning and optimum development of natural resources, activities clearly within purview of the state. Proper use of these facilities in support of exploration in a competitive, free-enterprise system should be based on achieving ultimate benefits to the public, but, in general, should not entail competition with commercial laboratories.

Jan 1, 1982

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 G. T. Adel

This paper presents the highlights of a general-purpose simulator for mineral processing plants that is being developed at the University of California, Berkeley. The state-of-the-art structured design methodogy based on a top-down design approach is being used to construct a modular and completely structured simulator such that it is independent of the unit operation models and routines. The simulator provides for two levels of operation allowing prebuilt or user-built simulation flowsheets depending upon the degree of computer knowledge expected of the user. The philosophy used in designing this simulator appears to be unique in the field of flowsheet simulation. The paper also includes a mini-example demonstrating the usage of the simulator.

Jan 1, 1982

By Constance F. Acton

This paper presents an overview of -- the carbon-in-pulp process for gold and silver recovery as seen in the early 1980's. Traced back to its origins in the mid-nineteenth century the process development through the present state-of-the art is described. The process technology and equipment design are described in detail for the carbon-in-pulp process. A typical process flowsheet is given with a description of appropriate design criteria. Technical advantages and disadvantages as compared to the traditional countercurrent decantation process are discussed including some illustrative comparisons of capital and operating costs. A summary of currently operating carbon-in- pulp plants throughout the world is given with an insight into future developments in this industry.

Jan 1, 1982

By Michael Hood

Subsidence profiles above two adjacent panels in Illinois are compared with pro- files predicting subsidence behavior obtained using the (i) National Coal Board (NCB, 1975) method, (ii) the profile function method and (iii) the influence function method. The NCB method predicts the maximum subsidence values at the center of the troughs accurately, but, the overall shapes of these predicted profiles do not match the profiles from the measured data well. Consistent values for the angles of draw were measured but these angles were different in the transverse (average value 43 degrees) and in the longitudinal (average value 17.5 degrees) directions. Time dependent subsidence effects are shown to be small but measurable. These displacements continue at a linear rate for at least a 12 month period. A comparison of measured horizontal distances, interpreted as horizontal strain, and the NCB predictions for strain shows that the peak measured strains are greater by a factor f about four than the predicted strains. The relationship between surface curvature and strain is investigated. Problems associated with calculation of surface curvatures from vertical displacement data are highlighted and a recommendation is made for future studies to consider direct measurement of this parameter. Surface curvatures above a moving face are found to be about three times less than the curvatures at the stationary end of the panel. On the other hand, a hyperbolic tangent profile function is shown to serve as an accurate predictive tool for subsidence behavior in two adjacent longwall panels at Old Ben Number 24 mine in Illinois. This function predicts not only the vertical displacements but also the surface curvatures above both panels. Influence functions are shown to be more problematic, although potentially more flexible, in their application than the profile functions.

Jan 1, 1982

By Ronald R. Bauer

Wollastonite is a little known industrial mineral of the silicate group. The largest and purest known deposit in the free world is being mined and beneficiated in Millsboro, New York by NYCO, A Division of Processed Minerals Incorporated. The associated garnet and diopside are separated from the wollastonite using dry high intensity magnetic separators. Beneficiated wollastonite is used extensively as an additive and filler in the ceramic, paint, refractory, and wallboard industries. Wollastonite acts as a functional filler due to its unique chemistry and physical acicularity.

Jan 1, 1982

By Robert L. Sundeen, Richard V. Wenberg

INTRODUCTION Shotcrete can be defined as spray-applied concrete with a maximum size aggregate in excess of 9.5 mm (0.375 in.). The concept evolved from guniting, which uses smaller aggregate and generally produces concrete with poorer physical properties. The larger aggregate used in shotcrete requires less water in the mix, while offering larger bonding surfaces, thus imparting greater tensile and compressive strengths. TYPES OF MACHINES Two basic designs of shotcrete machines in general use are the wet or prehydrated zero slump and low slump type, and the dry or nozzle hydrated machines. Fig. 1 shows a typical zero-slump machine in a con¬figuration commonly used for small-volume require¬ments. The prehydrated zero-slump machine mixes exactly proportioned batches in a sealed pressurized chamber. "Slugs" of hydrated shotcrete mix and com¬pressed air are sent through a distribution hose to the nozzle where the wet mix emerges as a spray. The spray is directed onto the rock surface by the nozzle operator. The low-slump machine is similar, except that a thick slurry mix of aggregate is delivered throughout the distribution hose to the nozzle. Fig. 2 shows a typical low-slump shotcrete machine used in moderate-volume applications. The nozzle-hydrated shotcrete machine accepts a premixed "dry" (4 to 6% moisture) batch of aggregate and cement. A constant feed of this mix is directed by compressed air through the distribution hose to the nozzle. At the nozzle, water for hydration is introduced into the emerging spray. The hydrated mixture is di¬rected onto the rock surface by the nozzle operator. Fig. 3 shows a typical dry-mix machine. The major difference between the two basic machine designs is the consistency of the shotcrete quality. Pre¬hydrated machines have high degrees of consistency and produce high-strength concrete, resulting from their ability to produce exact mixtures of zero or low-slump concrete within the machine. Nozzle-hydrated machines tend to produce poorer quality concrete because of the difficulty in controlling the water content at the nozzle. Water input essentially is an art. As such it is not precise, and the shotcrete produced is of variable quality. MATERIALS HANDLING Materials handling presents a problem to either type of shotcrete machine. At times considerable quantities

Jan 1, 1982

By David A. Campbell

There has been a long history of ironmaking on the North East coast of England centered around the River Tees. The industry developed here because of the availability of good coking coals to the North in Co. Durham, and the discovery in 1851 of ironstone in the nearby Cleveland Hills, South of the .River. Cleveland ironstone is relatively poor in quality, with an iron content not exceeding h% and high proportion of sulphur and phosphorous. The iron exists in chemical association with alumina and silica, and physical beneficiation processes involve a serious loss in iron unless the upgrading is limited, when the economics of the process become doubtful. Over the years, richer foreign ores have replaced the local ironstone, and the last operating mine was closed in 1964.

Jan 1, 1982

By Dave G. Osborne

With the growing exportation of coal from North America to the Pacific Rim countries and Europe, has come an increasing awareness of the need to determine economically optimized amounts of incombustibles (ash and moisture contents) for the shipped product. Despite the fact that demand bas caused a steady improvement in selling price, particularly for thermal coals, growing production and transportation costs have compelled producers to keep a sharp eye open for ways and means of minimizing the latter since this probably represents the greatest single potential for making savings. Consequently the Coal Industry has become acutely aware of the economic ramifications of eliminating moisture in particular, thereby producing further saving in the cost of transportation of product from the mine to the consumer. There is, however, a lower limit to which dewatering or drying can be taken and this relates to dustiness which is a function of the characteristics of the coal and the methods used for handling and storage during trans-shipment. Overdrying can lead to serious dust losses and difficulties in handling particularly if the coal contains a large proportion of fines. The ways and means for achieving this "happy medium" of optimization in moisture removal are seemingly undergoing gradual modification largely assisted by the following factors. a) Improvement in effectiveness of coal dewatering and drying equipment. b) Improvement in operational availability and utilization of the equipment. c) The development of surfactants, polymer flocculants and other agents which in many cases can be almost tailor made to suit specific applications.

Jan 1, 1982

By P. Diehl

The increasing public interest in a useful framework for an assessment and classification of mineral resources (1) is documented by the recently published international classification scheme of the United Nations (1979) and the revision of the classification system of USBM/USGS (2) (1980). Moreover, various classification systems are currently discussed by the mining industry. In spite of major improvements in respect of the concepts, definitions, and taxonomy of ore classification, it has been recognized that the recent practice to classify mineral resources does not create truly reproductible results, i.e. an assessment to the various categories which is free from personal bias or corporate interests. Following are geostatistical methods proposed for a standardization and quantification of various terms in ore classification. After critical review of present practices, we will establish a precise link between the classical terms such as "measured", "indicated", and "inferred" reserves and the geostatistical concept of estimation errors. Moreover, we will demonstrate, how the basic criteria of economic feasibility, geologic assurance and recovery can be taken into account by geostatistics. A specific algorithm will be proposed to block out and to classify reserves according to the degree of geologic assurance. . We will also show computerized methods which perform complete classification procedures, and will also exemplify them with typical case studies.

Jan 1, 1982

By Terry W. Engelhart

"A computer-based approach to online supervisory control and operations management of bulk materials handling terminals". This paper describes an integrated computer facility supporting the functions and responsibilities of shift operations, scheduling, maintenance and management using real-time interactive techniques. The functions and responsibilities as related to coal export terminals are specifically addressed, although these functions would be equally applicable to other terminal configurations and other commodities. A discussion of the resulting benefits are presented: - throughput efficiency - long-term planning and scheduling, using simulation - operations monitoring and protection - demurrage/despatch optimization - preventative maintenance - operational productivity

Jan 1, 1982

By B. Yarar

Samples of coal from an adit in the Fording River District of British Columbia showed hydrophilic properties between depths of 0-24.5 meters and was naturally hydrphobic at higher depths. Laboratory experiments indicated that weathered samples (i.e., those mined at 0-24.5m) could be upgraded by froth flotation at pH<2 following slimes removal. The i.e.p. of samples varied regularly between pH=1.6-4.1, the weathered fractions being at the lower end of the pH scale. These values followed a pattern comparable with the flotation behavior of samples. A patchwise structure for coal surfaces determining the resultant disjoining pressure, which in turn determines flotation-non flotation relationships is postulated.

Jan 1, 1982

By Y. J. Wang

Optimal sizing of a ventilation shaft is a classical subject in mine ventilation economics.. The shaft size is optimized by minimizing the sum of shaft-construction and ventilation-power costs, or owning and operating costs. The length and air quantity of the shaft are held constant, and the location of the shaft, affecting its depth, and construction and ventilation power costs, is not considered (Hartman, 1961; Lambrechts, 1974; Wang et al., 1978). This paper proposes an approach by which the location and size for a new pair of shafts (intake and return shafts) are optimized for a coal-mine ventilation network. Specifically, the problem is to determine the location and size for a new pair of shafts, and the distribution of air between the new and old shafts so that the combined cost of shaft construction and ventilation power is minimized. The shaft length or depth corresponds to the height of overburden at any shaft location. To simplify the problem, the shaft locations will be limited to the line along the main or submain entries, and the connection entries at the shaft bottom neglected. In order to calculate the ventilation-power cost, the air power of the system should be determined. This in turn requires the calculations of the air quantity and pressure of each fan in the network, satisfying Kirchhoff&apos;s laws and Atkinson equation. After obtaining the air power, the cost calculation is straightforward. In the following, we shall assume that the resistances of all airways, and the quantity for each mining section, which is controlled by adjusting the air regulator, are all known. The resistances for the shafts are determined from the estimated friction factors, and shaft dimensions and lengths. Furthermore, it will be assumed that the fans can be selected to meet the duties called for by any network solutions.

Jan 1, 1982

By K. Ueda

Lead concentrates from Kuroko (black ore) mines, which contain Cu 4 ? 7%, Pb 55%, Zn 3 - 7%, As 0.5 - 0.9% are treated by hydrometallurgical process in order to smelt the product by electric furnace smelting together with other lead resouces from Kuroko. The hydrometallurgical pretreatment involves sulfuric acid leaching at first stage, ferric leaching at second stage. This paper describes the short history of lead recovery from Kuroko and the performance of laboratory test, pilot plant and commercial operation.

Jan 1, 1982

By Duk-Won Park

A new gage for monitoring fractures in underground openings was successfully developed utilizing the reflection photo-elastic technique. The gage may be used to detect crack initiation in mine openings or tunnels so that remedial action may be taken to prevent further deterioration. Fringes concentrated around the cracks were spread by the use of rubber backing. Crack growth as little as 0.003 inch up to as much as 0.1 inch width were detected in this study. Variation of gage materials will provide alternative ranges. The feasibility of the method was tested and calibration was made in the laboratory study. Field study successfully demonstrated the capability of applications in detecting mine roof crack development.

Jan 1, 1982

By Christopher M. St. John, Michael P. Hardy

INTRODUCTION Geotechnical models, particularly those based on the finite element method, have been available to aid in en¬gineering design of underground mining excavations for over ten years. Despite this fact there are remarkably few cases of their use in mine design documented in the literature. It is therefore to be anticipated that many potential users of these models are relatively unaware of their capabilities and limitations and also of the form and detail of geotechnical data needed for their success¬ful application. This subsection attempts to address this problem by discussing the different types of numerical models now available and by noting how some of them have been used to study a variety of problems associated with underground mining. The subsection concludes by discussing the applica¬tion of the displacement discontinuity method to the de¬sign of possible mining systems for the copper-nickel deposits of northeastern Minnesota. The object of the analyses, which were nonsite specific, was to determine the significance of geotechnical parameters, such as ini¬tial stress and rock structure, on the stability of under¬ground excavations and hence to provide guidance for future geotechnical investigation. NUMERICAL MODELS In the geomechanical design of underground mining openings, use is made of numerical models that repre¬sent or simulate the large-scale mechanical behavior of rock. There has been less interest in analysis involving fluid flow and heat transfer, but with increasing interest in such areas as in-situ retorting and solution mining it is likely that there will be a growing need for numerical models embracing and coupling all three physical proc¬esses. However, the emphasis in this subsection will be on mechanical behavior. Models which simulate such behavior will be divided into two groups: continuum models and discontinuum models. These will be dis¬cussed in turn in order that some insight into alternative solution strategies and their merits may be gained. Continuum Models Almost all geomechanical numerical models must be classed as continuum models even though particular computer codes incorporate special provisions for rep¬resenting discontinuities such as faults, bedding planes, or joints. They are continuum models because they pro¬vide solutions for cases where material behavior is governed by the differential equations of continuum mechanics. Two basic solution strategies for such equa¬tions may be identified immediately: the differential ap¬proach and the integral approach. In the differential approach a means of approximating the differential equations over the entire region of interest is sought. In the integral approach use is made of fundamental solutions from continuum mechanics, and these are used to construct a solution to the whole problem, making approximations only on the boundaries of the region of interest. The several differential and integral methods are identified in Table 1. Differential Methods: Problems in continuum me¬chanics involve the solution of three types of partial differential equations. Two of these govern the behavior in so-called initial value problems, in which variables change both in time and in space. Examples of such problems include nonsteady heat transfer and fluid flow, and stress wave propagation. The last type of partial differential equation governs the behavior in boundary value problems. In these, variation is in space but not in time. Solution of initial value problems may be achieved in two significantly different methods: implicit and ex¬plicit. The differences between these two methods will be illustrated by considering a very simple initial value problem, that of one-dimensional heat diffusion. The equation governing this process may be written as: [ ] where T is temperature, K is thermal conductivity, p is density, c is specific heat, t is time, and x is the spacial coordinate. In finite difference form this equation might be written as: [ ] where the superscripts refer to the time and the sub¬scripts to the spacial location. Several solution strategies for this equation have been used. Two important ones may be illustrated very simply by discussing the signifi¬cance of the superscript * on the right-hand side of the equation. If i + 1 is substituted for * then the second derivative of the temperature with respect to distance is evaluated at the end of the next time step (using tem¬peratures not already known). Such an approach leads to a set of equations involving unknown temperatures and a solution procedure which is referred to as being implicit. The important characteristic of the implicit procedure is that it leads to a set of equations that must be solved for each time at which the temperature dis¬tribution is required. If instead of substituting (i + 1) for the superscript *, i is substituted, the following equation is obtained: [ ] {In this case the new temperature is defined in terms of an already known temperature distribution. The solution procedure is now known as explicit and has the impor¬tant characteristic that there are no equations that have to be stored or solved. A practical advantage of this

Jan 1, 1982

By F. L. Wilke

Truck haulage costs form a major part of the total mining costs in hard rock open pit, mining operations. In order to minimize these costs one has to - determine the optimal truck fleet (type, size and number of trucks), and. simultaneously - determine the optimal organization of running the haulage system, i.e. especially select the optimal dispatching strategy in order to ensure maximal utilization of the trucks, and simultaneously - establish a reliable maintenance and servicing system to guarantee maximal availability.

Jan 1, 1982

By Donald O. Rausch, William E. Hawkes, Odin Jr. Strandberg

INTRODUCTION Kennecott Copper Corp.&apos;s Tintic Div. Burgin mine is a 725 t/d (800 stpd) lead-zinc producer, located in the east Tintic mining district, 5.6 km (3.5 miles) east of Eureka, and 130 km (80 miles) south of Salt Lake City, UT. The Burgin mining complex includes three shafts, shop, dry, and office facilities, and a 454 t/d (500 stpd) flotation concentrator. Production and ser¬vice entry to the mine is through a 4.26-m (14-ft) diam concrete-lined production shaft, and a three-compart¬ment rectangular shaft. A connection with the Apex No. 2 shaft and the Burgin mine provides a second exhaust airway. Bear Creek Mining Co., Kennecott&apos;s exploration sub¬sidiary, began exploration of the Burgin mine area in 1957 with the sinking of the 335 m (1100 ft) deep No. 1 shaft. A station was cut at the 320-m (1050-ft) level of the shaft and extensive drifting, diamond drilling, and some winzing and raising confirmed the presence of a major lead-zinc ore body, the main portion of which oc¬curred below a hot saline water table. Surface facilities and the 4.26-m (14-ft) diam concrete-lined production shaft were completed in 1966. Ore production has been steadily increased to 725 t/ d (800 stpd) since 1966. The collar elevation of the No. 2 production shaft is 1731 m (5678 ft) above sea level. The mine&apos;s three levels, the 1050, 1200, and 1300 are at elevations of 1414.6, 1356.4, and 1338.1 m (4587, 4450, and 4390 ft), respectively. The approximate elevation of the hot saline water table is 1382.6 m (4536 ft). The earliest production on the east Tintic mining district was from the Eureka Lilly mine in 1909. Later, in 1916, the Tintic Standard mine was discovered, and pro¬duced ores with a gross value of $80,000,000 through 1956. The Eureka Standard mine located 1.6 km (1 mile) southwest of the Burgin mine produced ore worth $12,000,000 from fissures in the Tintic quartzite along the Eureka Standard fault. Ores produced from the east Tintic mining district through 1971 are valued at $151.3 million. The production from the entire Tintic district is approximately $500 million. GEOLOGY The Burgin ore deposit occurs as replacement bodies in limestone, dolomite, and shale. Numerous fault struc¬tures and rock formations make the mine geology com¬plex. The mineralization is strongly associated with sev¬eral major geological features, which include the Ophir formation, consisting of limestone, dolomite, and shale as the favorable host rock, and the structural influence of the east Tintic thrust, the Eureka Standard and sev¬eral other significant faults. Fig. 1 depicts the Burgin 1300 level workings, a generalized configuration of the ore body, and the main geological features. Covering the quartzite, limestone, dolomite, and shale facies enclosing the Burgin ore deposit are 183 to 274 m (600 to 900 ft) of andesite. The Burgin ore consists of massive sulfide and oxide lead-zinc ores with significant silver and cadmium val¬ues. The main ore minerals are galena, cerrusite, and spalerite. Extensive manganese oxide deposits are associated with the lead-zinc ore occurrence. A limited

Jan 1, 1982

By Hans Hamrin, Bruce Kennedy

Once an ore body has been probed and outlined, and sufficient information has been collected to warrant further analysis, the important process of selecting the most appropriate method or methods for mining can begin. At this stage, the selection is preliminary only, serving as the basis for a project layout and feasibility study. Later, it may be found necessary to revise the details, but the basic principles for the ore extraction should remain a part of the final layout. With respect to the basic principles employed, relatively few mining methods are used today. For the purposes of this presentation, the principles are classified as follows: 1) For naturally supported stopes, the available methods are room-and-pillar mining and sublevel stoping. 2) For artificially supported stopes, the available methods are shrinkage stoping, cut-and-fill mining, square-set mining, and longwall mining. 3) For caving methods, the available techniques are sublevel caving and block caving. Due to the uniqueness of each ore deposit, the variations on each of these methods are nearly limitless. It is impossible to include even the major variations in this chapter; the goal of this chapter is to present the principles as a guide to the mining engineer in adapting the general methods to specific needs. The selection process described herein is intended to supply the techniques by which the candidate methods available for a given ore body can be reduced to one or two feasible approaches. The feasible approaches then can be evaluated in detail, and the particular modifications can be investigated. In describing costs and economic factors, all values herein are expressed in terms of 1978 US dollars.

Jan 1, 1982