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By Kadri Dagdelen

Design of the ultimate pit limit to direct long range mine planning has been a celebrated problem in the mining industry, and in mining academia, since 1965. Although many papers have been presented on this topic, even today, there is not a single solution technique which is widely accepted and considered universally best. Many philosophical differences exist among the pit planners concerning the use of different methods and it appears that resistance is very strong towards accepting other techniques if you are a believer in one. Whatever the case might be, the pit parameterization technique for the ultimate pit limit design is the latest and takes a different approach than its predecessors. There are already a number of papers which discuss the concepts involved (Francois-Bongarcon, 1976; 1978; 1980; 1981); however, the aim of this paper is to fill in some of the gaps which are thought to be missing in the previous discussions and, in addition, to present the recent research done on the subject at the Centre de Geostatistique de l?Ecole de Mines de Paris in Fontainbleau, France. This paper is divided into two parts. The first part will discuss practical aspects and interpretations of pit parameterization concepts with- out using mathematical or theoretical arguments; the reader who is interested in the theory should refer to Francois-Bongarcon (1978). The second part will present an approach to obtain a complete double parameterization of reserves using an inexpensive (computer costs) single parameterization algorithm. A case study of this approach is also included.

Jan 1, 1982

By Jack V. Hill

A unique type of grinding mill liner has been developed by Waagner-Biro A.G. of Austria. This liner system, termed an angular spiral lining (.ESL), was, successfully tested by Magma Copper Company's San Manuel concentrator in both their 10 Ft x 10 Ft (3.05m x 3.05m) and 12 Ft, 6 In x 14 Ft (3.81m x 4.27m) grate discharge wet grinding ball mills. Magma Copper's savings in power and grinding balls justified conversion of more than half of their twenty-six ball mills to the ESL system.1,2 The ASL system was provided by the Capitol Castings Division of Midland Ross Corporation, the exclusive licensee for such conversions. Capitol Castings proposed the following as expected benefits from installation of an ASL system in an 18 Ft x 21 Ft (5.49m x 6.40m) overflow discharge wet grinding ball mill at the Pinto Valley Operation of Cities Service Company: "(a) Reduction of power consumption per ton of ore milled = 12%, (b) Increased throughput = 6%, (c) Reduction of ball consumption = 6%."

Jan 1, 1982

By Roger M. Buckeridge, Charles D. Reese, Silvio M. Martino, William T. Carey, Ansel H. Graham

FUNCTIONS OF RAIL HAULAGE SYSTEMS The principal application of rail haulage in under¬ground mining is in the movement of ore or coal pro¬duction from near the mining areas out to a processing plant. In some cases, such production haulage may end at a shaft pocket or slope belt conveyor transfer station, depending on the type of mine and the geo¬logical placement of the mineral being mined. Main-Line Haulage The basic reasons for choosing main-line rail haul¬age are high-tonnage production per shift and long haulage distances. The choice between rail and other systems, such as belt conveyors, is not easy and requires an integrated study of many factors: tonnage, distance, grade, product size, number of producing areas, ore reserve, safety, capital cost, operational and maintenance cost, and the mining system. Projecting a main-line rail haulage system for com¬parison with other systems requires a choice of many variables which are interrelated. The project must start with consideration of the mining system, tonnages per day, distances and grades, and loading and dumping conditions. Rail mine cars of different sizes and types are studied. High tonnage demands consideration of wider track gages. A suggested minimum for metal mines is 914.4 mm (36 in.); for coal mines, 1066.8 mm (42 in.). The choice of the train size should be governed by tonnage per shift, physical factors of loading and dumping, cycle time, and the number of trains that can be operated and safely controlled. After the size of the train is determined, the type, size, and specifi¬cations of the locomotive can be selected. The loco¬motive may be trolley, battery, or diesel; selection guides are discussed elsewhere in this chapter. The advantages of rail main-line haulage are high tonnage capacity, flexibility, reliability, safety, low operational and maintenance cost, and dual usage of the system (supply haulage, for example). The dis¬advantages are high capital investment, high installation cost, the requirement for more maintenance facilities, and restriction to certain ore reserves and mine pro¬jections. Supply Haulage Supply haulage is a basic requirement of any mine, but the type and volume of supplies vary with the type of mineral being mined, the mining methods, and the size of the mine. A mine developed for rail main¬line haulage gains supply haulage through the addition of a secondary track system to extend the main line closer to the producing areas. Additional haulage equipment, generally smaller, is also added. It is neces¬sary to design into the rail system more switch bypass points and train control systems. The track system standards demanded by rail main-line haulage con¬tribute to the reliability and efficiency of the supply haulage function, whereas a supply track system to supplement other haulage systems often suffers from poorer track and installation standards. When belt conveyors are chosen for production haulage, a rail supply haulage system is often installed rather than using rubber-tired vehicles for supply haul¬age. The conditions of the seam or ore body create these items for consideration: haulage grade of more than 6% favors rubber-tired vehicles; longer haulage distances favor rail; difficulty in maintaining a smooth roadway favors rail; and movement beyond the end of a rail system generally requires rubber-tired vehicles, making a single system rather than a dual system attractive. Although operation of a supply haulage system may be concentrated on a single shift of a multiple-shift mining operation, proper planning and supervision are required if the haulage is to be efficient. Often the newer personnel at a mine are assigned to supply haul¬age which may add to the difficulty of obtaining maxi¬mum efficiency. Supply haulage by rail offers the advantages of high reliability, lower maintenance, and high efficiency over long distances. The disadvantages are less flexibility, high installation costs, and rubber-tired vehicles for use beyond the inby end of track often required. Personnel Movement The efficient and comfortable movement of the miner to and from his work area has a significant effect on his daily output. Minimizing man-trip time also becomes more and more important as hourly wages increase. Personnel are generally moved over a rail system by one of three methods: a specialized personnel carrier to which a certain number of men who work in a certain area are assigned; a combination vehicle, used to carry personnel but also capable of being used as a light supply/utility locomotive during a working shift; and man-trip cars towed by a supply locomotive. The first and second methods are used in coal mines where long distances make speed and comfort important. The third method is more common in metal mines. Equipment Movement As mining equipment technology has advanced, many mining machines and their subcomponents have grown in size. This trend requires careful planning for adding specialized carriers to handle the movement of larger equipment. Continuous miners, longwall equipment, raise borers, and ventilation equipment are examples of machines which may be difficult to move and which require special handling. The selection of the weight, speed, and operating characteristics of supply locomotives used to move equipment is important. Larger mines may often justify a special group of loco¬motives for this application.

Jan 1, 1982

By K. -A. Björkstedt

INTRODUCTION Strassa lies in the central part of Bergslagen, a tradi¬tional mining district, on the eastern side of the Stora Valley at an elevation of about 200 m above sea level. A railway siding runs between the mine and the Stora railway station from which there are railway connections to the shipping port and iron and steel works in Oxelosund, about 224 km away. The distance to the provin¬cial capital Orebro is about 60 km. The climate is typi¬cal for this part of central Sweden and is illustrated by the diagram of monthly precipitation and temperatures for the years 1968-1975 (Fig. 1). HISTORY There is no certain information as to when the Strassa mine was first worked, but it is known from sur¬viving accounts of mine inspectors that there were smelt¬ing works in operation in nearby villages in the 12th century. An example is the Gusselhytta ore smelting works, 10 km south of Strassa, which dates from this period. Around the year 1540 there were two smelting works in Strassa, the Upper Karberg and Lower Karberg works. Ore for these smelters was probably taken from Strassa and from the adjacent Blanka mine. In the year 1624 Strassa is mentioned by the painter Jons Nils Krook in an account of the iron mines in the Linde mining district (Linde Bergslags Jarngruvor). Several mines were listed in the area, the deepest being about 30 m. An impressive power installation is mentioned in 1639, including a piston system of lashed poles for transmit¬ting power from the Stora River to the Strassa fields. Its length was 2670 m. Common ground comprising about 20.2 km2 (5000 acres) of forest was allocated in 1689 for the furtherance of mining operations. Until the beginning of this century only the rich cen¬tral parts of the ore body were mined and these yielded, after handpicking, lump ore suitable for smelter feed. An example of the ore grades from these early times is an analysis of ore from the "Big Mine" (Storgruvan) from the year 1873: 48.5% Fe, 0.008% P, and 0.06% S. This same year a total of about 18 000 t was ex¬tracted from the Strassa mine. OWNERSHIP The mine was owned and run until 1874 by a min¬ing association made up of 119 so-called "bergsman," who were homesteaders often engaged in agriculture and timber-cutting as well. In that year the Strossa Grufvebolag (Mining Co.) was founded. In 1906 it was con¬verted into a joint stock company, the Strossa Gruveaktiebolag. This was acquired in 1907 by Metallurgiska AB for the implementation of Gustav Grondal's beneficiating and briquetting methods, for which the Strassa ore was well suited. The same year saw the completion of a new ore dressing plant with an annual production of 46 000 t of ore concentrate. In 1911 the mine passed to new hands, and in 1913 it was purchased by an Austrian company. Extensive new installations were made and in 1915 a new dressing and briquetting plant was completed with twice the capacity of the old one. In 1917 the Strassa mine was acquired by Granges. Be¬cause of unfavorable business trends and technical diffi¬culties, mining operations were brought to a close in 1923. Pumping kept the mine free of water until 1933 but it was completely filled ten years later. Up to 1950 the surface buildings and installations remained intact but the large dressing and briquetting plant burned to the ground in that year. Today only the machine shop re¬mains from this earlier period of operation, now housing parts of the Mineral Processing Laboratory. The decision to take up mining operations again was made in 1955 and construction work began the follow¬ing year. Of the old installation, only the "southern shaft" could be used for some development drifting after it had been completed with a new headframe. Other¬wise, all the buildings and installations required for the operations had to be rebuilt. New installations ready by 1960 were office and personnel facilities, a new shaft and headframe, a sorting and concentrating plant, a macadam plant, settling basins, pump stations, and a railway and yard with transport equipment. The instal¬lation was completed with two plants

Jan 1, 1982

By A. D. S. Gillies

The effectiveness of coal mine face ventilation and the ability of air flow to maintain satisfactory methane and respirable dust levels in the working face area has been studied through the use of scale modeling techniques and underground testing procedures. In particular, consideration has been given to examination of the effects of operation of a high air flow machine mounted dust scrubber unit on face air patterns. Results obtained through a cooperative study with the Peabody Coal Company demonstrate that new mining cycles can be considered which increase production efficiency while maintaining satisfactory ventilation conditions.

Jan 1, 1982

By John L. Halls

INTRODUCTION In this section, it is assumed that the reader is an engineer with sufficient training in underground mining work, preferably with budgetary control, to understand the mining methods outlined in other sections of this handbook. The mining-method descriptions are not repeated herein; only sufficient information is provided to enable the reader to comprehend the development of the costs presented. All costs are in 1977 US dollars, except where explicitly stated otherwise. CAPITAL INVESTMENT The capital investment in the underground mining portion of any mining property includes the following: 1) The development of original access involving adits, shafts, declines, etc. It includes the necessary equipment for transporting men, materials, and rock. 2) The purchase of mine equipment including loaders, underground haulage equipment, fans, pumps, electrical distribution, surface facilities associated with the underground mine, etc. 3) Sufficient preproduction mine development to assure steadily increasing production after startup. De- pending upon the mining method used, the preproduction development should provide a supply of three months to one year of ore that is developed and is ready to be stoped. Underground development also includes excavations other than shaft stations, loading pockets, sumps, etc., which usually are associated with the means of access and are excavated by the contractor. These excavations could include a crusher cavern, an ore- storage pocket, and underground workshops that may or may not be included by the shaft-sinking contractor. Portions of these could be delayed until after production begins, but it is preferable to complete them during the development phase, so there is no interference with the production buildup. 4) The staff and labor costs for all phases from inception through final production. 5) The provision of mine-surface equipment and facilities strictly applicable to the underground portion of the operation. The required facilities include a changehouse with a lamp room, a first-aid station, compressors, an electrical substation, workshops, a ware- house, water pipelines to the concentrator, water storage tanks, explosive storage, etc. A concrete-batch plant, a sand-fill plant, and boilers for heating may be required. 6) Other direct and indirect costs. 7) Engineering, design, and procurement costs. 8) Contingency allocations. 9) Contractor's fees, startup costs, and owner's costs. Original Access As part of a study for the US Bureau of Mines, Dravo Corp. conducted an extensive literature research survey covering underground operations in the United States between 1963 and 1973. Table 1 is compiled from the section of that study pertaining to the types of access to underground mines. Table 1 does not include distances attributable to the deepening of existing shafts, nor does it include shafts bored by raise drills or downhole machines that need existing underground workings for the disposal of the cuttings. During the period under review in the report ( 1963 through 1973). concrete-lined circular vertical shafts were the predominant means of providing the first access to a new underground mine; these shafts accounted for 80% of the original access in terms of the number of accesses completed, and 75% in terms of the total verti- cal linear distances. The inside diameters of the conventional circular shafts ranged from 3.7 to 7.0 m (12 to 23 ft). Most (72% ) of these conventional shafts were for production,

Jan 1, 1982

By Mehmet Guney

Acceptable working environments in production areas and service openings of Mindola Mine on the Zambian Copperbelt are currently maintained with an air volume of 1000 cu m/s and 5500 kW of installed refrigeration plant. It is planned to open up additional deeper areas of the mine for future production. Temperature problems, due to the increased heat generation from various underground sources are recognized as potential hinderances to efficient production. Findings of research work indicate that the present total heat output of 58,000 kW will inevitably increase by the second five-year period of the deepening program to 52,900 kW and by the fifth five-year period to 63,200 k\V. Estimated WB and DB temperatures, representing the total heat generation in stope faces and development headings during these same periods are predicted to increase to 53.8/40.2 °C and 35.5/42.1 °C respectively. It is assessed that the nominal refrigeration capacity needs to be progressively increased to 9,000 k\V This paper deals with the thermal problems in Mindola Mine, current heat output calculations, future prediction of temperature and heat load in various working areas of the mine with the application of computer programs and interpretation of data in order to find the effective solution for long-term planning of ventilation and air-conditioning.

Jan 1, 1982

By G. W. Johnston

INTRODUCTION For many years mining at the Stobie mine of the Inco Metals Co. consisted mainly of fill-type mining for higher grade ores and block caving or blasthole open stope mining for lower grade ores. Open stope mining, although quite favorable in the initial stoping phase, has always presented problems in the pillar recovery phase, especially at depth. In 1966, sublevel caving was introduced. This method is ideally suited for low-grade ore and can be applied very successfully to intermediate or higher grade ores providing proper controls are enforced. Sublevel caving has three distinct advantages: it is highly adaptable to mechanization; it recovers the ore in one operation; and it provides adequate ground sup¬port and safety regardless of mining depth. This method does, however, make extra demands, particularly for new types of supporting services. The constant presence of rock makes quality control of the grade of the ore being extracted essential. Increased de¬velopment, different types of ventilation, roadways, and maintenance are only a few of the many changes re¬quired. The care and maintenance of the fleets of diesel equipment and the modification of equipment is a con¬stant challenge. CHOICE OF SUBLEVEL CAVING The question arises, why mine by sublevel caving? Fig. 1 shows a typical problem encountered by the open stope or blasthole mining method. Near the surface there is little or no ground support problem. As mining progresses to an intermediate depth, however, ground failure starts to occur in the supporting pillars. If open mining is carried beyond this point, the critical depth is soon reached, accompanied by hanging wall caving, ex¬tensive pillar failure, and the danger of losing control of the mining cycle. The rock caved from the hanging wall mixes with and dilutes the ore being extracted from the blasthole stopes. Ore losses occur when pillar failure pre¬vents drilling and development of part of the ore. The interruption to the mining cycle causes serious curtail¬ment in production until ground failure is controlled and the mining method changed to adapt to the new conditions. Having recognized the limitations of open stope min¬ing, alternatives were sought. The first alternative (Fig. 2) was to fill the open stopes with sand fill be¬fore extensive pillar deterioration occurred, recover the pillar ore by fill mining, and then convert to total fill mining below. If the ore grade is high enough it can support the costly fill method. Recovery will be at a much slower rate and will therefore extend over a longer period of time. On the other hand, sublevel caving gives complete primary mining recovery in a short time and is very adaptable to mechanization, which is now a necessity to combat spiralling labor costs and give greater produc¬tivity per man. Sublevel caving is ideally suited as a mining method for low-grade ore bodies which eventually reach a criti¬cal depth for open stope mining. DEVELOPMENT The development phase comes first and of course its main purpose is to gain access to the ore and provide a route for ore removal from the mine. Initially a 20% inclined service ramp is driven in the footwall rock (Figs. 3, 4, and 5). It is preferable to start at the top of a mining block and drive the ramp down¬ grade so gravity can assist in mucking the blasted de¬velopment round, and also access can be gained more quickly to the initial production area. The ramp nor¬mally turns at 3.1 rad (180°) about every 152.4 m (500 ft) to keep it reasonably close to the rock disposal

Jan 1, 1982

By Daryl E. Radcliffe, Raymond M. Stateham

The Bureau of Mines, under a Cooperative Agreement with the Bear Coal Company, Somerset, Colorado, conducted a large scale evaluation of the effect of time lapse on roof stability. Time lapse in this study is the interval between exposure and permanent support of the mine roof. The study indicates that time lapse is not a critical factor in roof stability provided permanent support is installed before catastrophic failure begins. Concurrently with the time lapse study, investigations were made of stress distribution around resin-grouted bolts, the regional stress field, and anomalous geologic conditions. The behavior of the resin-grouted bolts, at least in the Bear mine, appears to fit the Coulomb-Navier Theory. Geologic conditions are primary factors in trends of roof stability that extend across the mine.

Jan 1, 1982

By E. M. Van Eeckhout

The National Uranium Resource Evaluation Program (NURE) has produced a mass of geochemical data in the form of data files on tapes. This paper presents a methodology for utilizing those tapes in a simple, practical manner. An example is given of an analysis of the Butte quadrangle, Montana, tape obtained from the Los Alamos National Laboratory (LANL). The example details the sequence of splitting the data file, plotting histograms, cumulative frequency curves, concentration plots, and methods for evaluation. All programs are written in FORTRAN. The case study was performed using a mini- computer (DEC PDP 11/70) and standard plotting subroutines.

Jan 1, 1982

By Louis A. Panek

INTRODUCTION Block caving is distinguished from other types of mining in its requirements that the mineral body to be extracted must cave after a relatively large tabular open¬ing is created underneath it, and that the caving must be predictable and controllable for both operating and safety reasons. A mineral body that caves readily may present special problems with respect to ground support in the mining access openings, where stability, rather than failure, is the desired objective. The complete ex¬traction of a caving block, which may measure 50 to 100 m on a side and 100 to 300 m in height, tends to create a large void, resulting in significant changes in the ground stress system and a migration of rock from all directions toward the void; these events begin very early in the life of the mine and lead to additional spe¬cial problems with respect to control of ground move¬ments in the mine working areas, in the rock adjacent to the caving stope, and on the subsiding surface above the mine. Thus, the mechanics of an undercut-cave operation involve a broad range of geotechnical subject matter, including the determination of the lithostatic earth stresses, the structural properties of a large jointed rock mass, the flow properties of the fragments after caving, the stability of tunnel-like openings, and the stability and displacements of the rock structure (1 to 2 km in size) that surrounds the ore body. The objectives are achieved by in-place measurement of rock properties, laboratory tests, theoretical structural analyses of the related prob¬lems, measurements of structural behavior in and above the mine, and comparison of theoretical predictions with actual performance in order to interpret and refine the predictive capability. Undercut-cave mining (Julin, et al., 1973) differs from most other methods in that the caving is an essen¬tial and inherent part of the excavation process and in¬volves the movements of millions of cubic meters of rock over distances of hundreds of meters, as compared to the gradual 2-m settling of the overburden typical of full¬caving longwall coal mining. Undercut-cave mining may involve a rock mass on the scale of a 1-km diam open¬pit mine, but mining is several times deeper than that in the open-pit mine; furthermore, a large portion of the rock mass is in a failed, rather than a stable, state. Ground-control problems are inevitable in a situa¬tion where ground movements are occurring on such a large scale, and where the rock mass is expected con¬veniently to cave at one location but to be stable at an¬other. To devise effective responses to these ground¬control problems, one needs to know first of all what is happening, not merely at the point where the symptoms are observed, but more broadly with respect to the be¬havior of the entire mine rock-mass structure over the total area of extraction, including the effects on the sur¬face as well as those underground. Monitoring structural behavior on this scale would require a sizable group of people, if they attempted to measure all the significant mining-induced changes of ground displacement and pressure, even on a 100-m grid of points. The practical approach is to conduct the mea¬surements activity by a judicious sampling, making in¬termittent, concentrated efforts in selected areas of lim¬ited extent to study problems of immediate importance; simultaneously a minimum level of surveillance is main¬tained over the total structure to delineate long-term be¬havior trends in the major structure, which provide the baseline reference data that are so essential for interpre¬tive purposes. Mine-structure behavior depends not only on the mine extraction configuration and sequence, but on the rock structural properties and how they change from one location to another in the mine. Increasing attention is being given to determinations of rock strength and de¬formation by performing mechanical properties tests in place and in the laboratory, and by drawing inferences from the mapping of geologic structure in mine openings as they are driven and from the logging of cores ob¬tained from exploration drilling that may be performed primarily for delineating ore reserves. Monitoring of mine-structure behavior and inter¬preting it in terms of the growing excavations and the structural properties of the rocks have significant im¬pacts on designing the major undercut-cave operations of undercutting and drawing the ore, supporting the ac¬cess openings, and controlling surface subsidence. The purpose of this chapter is to call attention to the com¬mon ground-control problems encountered in undercut¬cave mining, to relate them to the mining process, and to explain the geomechanical approaches to analyzing these problems with emphasis on methods that are most readily available to the mine staff. UNDERCUTTING, CAVING, AND DRAWING A distinguishing characteristic of an undercut-caving method of extraction is the objective that the rock mass fails of its own weight in a predictable and controllable manner, after it is undercut by excavating a tabular¬shaped void, horizontal or inclined. Induced or forced caving refers to procedures that may be resorted to when the ore will not self-cave over the desired span. Block caving, panel caving, and mass caving are terms used to denote the particular sequence of undercutting and ex¬tracting the total area of the ore body on a single min¬ing level. The discussion herein is not specialized to any of these variations; the word block as used does not imply anything as to sequence. CREATING THE UNDERCUT A common undercutting method is to drive a series of drifts and blast the intervening pillars. Theoretical structural analysis and practical experience show that ground-support problems in the undercutting drifts and in underlying access openings as well (the same princi¬ple applies even if undercutting is done by belling out drawpoints from the slusher drifts) are minimized by retreat-mining across the undercut slot according to a sequence such that the growing undercut area retains a

Jan 1, 1982

By Peter I. Brooker

The semi-variogram, the basic tool of geostatistical estimation procedures, must itself be estimated from the available data. From a numerically simulated two dimensional deposit, with known variability expressed through its semi-variogram, the standard arithmetic mean estimator is used to calculate experimental semi-variograms. These are determined on a series of data sets obtained from the deposit with various sampling grids which cover different areas of the deposit and which have a wide range of available data. The orders of discrepancies between estimated and theoretical semi-variograms are determined in this idealized situation; and this provides a basis against which to assess experimental semi- variograms made on field data.

Jan 1, 1982

By A. L. Mular

Se revisan brevemente varios procedimientos para el cálculo de coste de capital; no obstante, se discute en preferencia el método del factor estimado. Se proporciona una tabla de costes frente a los datos de parámetros de equipo, de donde los costes pueden ser adaptados a partir de los valores actuales de los índices de costes M&S para minería y molienda. Las ecuaciones de costes son utiles para preparar los programas de computador que realizan cálculos de coste de capital. Se ilustra la utilización del metodo del factor estimado.

Jan 1, 1982

By R. O. Burt

The paper examines the role of gravity concentration in both an historical contest and within the modern mineral processing industry. It discusses the principles of the process on both a theoretical and practical level. The importance of ore mineralogy and its effect on the choice of the wide range of equipment available is discussed. The paper concludes with a discussion on the relative merits of gravity concentration to other forms of processing and seeks to answer the classical question - "Quo Vadis?"

Jan 1, 1982

By Stanislaw Siewierski

One of the major problems in the exploitation of deep copper ore deposits in the Legnicko-Glogowski Copper Field is ground control and, in particular, roof control of the headings. The depth of copper deposits ranges from 600 to 1200 m. The parameters of ore-bearing rocks and the surrounding rocks show considerable variance of strength properties along the mining thickness (Fig. 1). The roof of the openings exploitation headings consists of dolomite-calciferous rocks. The roof strata consist of layers, the thickness of which vary from several centimeters to 0.6 m (sometimes to 1.0 m). These strata differ in strength and are characterized by igneous and sedimentary discontinuities. On the basis of studies in structure, properties and geological mining conditions of the roof strata in Polish copper mines, the method of roof classification has been developed. (1) This method uses the stability index "L" of the roof, calculated from [ ] The value of "t" ranges from 0, l to 0,8. Based on the value of "L", four classes of the roof (Table 1) corresponding to different degrees of fall hazard have been developed. The mining of copper ore deposit is carried out by a single lift room-and-pillar mining system or a two-stage room-and-pillar system with roof caving. (3, 4). The single lift room-and-pillar mining system is most often used. It consists in working of the deposit of 2.5 to 4.5 m mining thickness, 150-350 m long mining line, consisting of parallel room faces 6 m wide with pillars 5.7 or 16 m wide. As the mining line advances, cross-cuts joining individual rooms are made, thus forming support pillars 3x5, 7x7, 5x16, and 7x16 m (Fig. 2). The two-stage system (Fig. 3) mines the deposit with rooms 6 m wide into blocks 25 x 35 x 25 x 45 m. During the second step these pillars are worked with rooms and cross-cuts into pillars about 3 x 4 to 4 x 5 m (retreat mining). The pillars are crushed just before the fall of the roof. The maximum opening of workings in the two-step system is about 10 m, i.e. one belt entry and one row of bearing pillars. In single lift room-and-pillar systems it is about 35 m., i.e. three belt entries and three rows of bearing pillars. The support of the headings consists of expansion bolts 1.6-1.8 m installed in a bolting pattern 1.1 x 1.5 m. In both systems the fall step is about 10 m, i.e. the width of a belt and a bearing pillar. The Indications of Rock Mass Pressure Individual Dog Heading Individual dog headings are set wide apart from each other in the unmined ore deposit area where there is no interaction of headings. The indication of rock mass pressure are, in such cases, limited to the deformations and damage of side walls. Radiometric studies of rock mass loosening showed that the maximum range of lifts in side-walls immediately after web was 2 m (4,5). In the course of time the cracked side rocks fall off, thus widening the heading. The morphology of side-walls is influenced by the location of the heading in relation to clump layer/tarry. This layer is characterized by a weak strength and "lubricating" properties. The location of the heading in

Jan 1, 1982

By Yoginder P. Chugh, M. Silverman

Premining investigations involve the determination of roof, floor, and coal seam characteristics to delineate areas of troublesome ground control prior to development of a mine. Data obtained can be useful in mine design, equipment selection, estimation of face productivity, development of ground control plans and assessing ground control costs, and in conducting feasibility studies and financial analyses. This paper discusses 1) the causative factors associated with ground instability, 2) how useful data can be obtained for these factors during premining investigations, and 3) analysis techniques to delineate probable areas of ground instability.

Jan 1, 1982

By R. W. Chadwick

Coal was discovered in South Africa in the mid 1830's. The earliest exploitation of modern times was probably in Natal where, in a letter dated 20 August 1842, a Captain Smith mentions that a wagonload of coal brought to Pietermaritzburg, was sold. Until the discovery of diamonds in 1867 and the Witwatersrand gold in 1886, coal mining satisfied a very small domestic demand. By 1889, at least four collieries were operating in the Middelburg-Witbank district, some 120 kilometres East of Johannesburg, and mining had star- ted at Vereeniging, about 50 kilometres to the South of Johannesburg. A map of South Africa is shown in Figure 1 and an enlargement of the main coal mining areas in Figure 2.

Jan 1, 1982

By Neil J. Dahmen, Howard J. Handewith

INTRODUCTION Tunneling machines were introduced into civil con¬struction in the 1800s, but their practical applications were not realized fully until the 1950s. As of 1977, tunneling machines were being used on most tunneling projects of substantial size. As in civil construction, the mining applications of tunneling machines have included long drives such as tunneling for long ore-body develop¬ment drifts, long haulage tunnels, longwall access drifts, and long crosscut developments. Tunneling machines ranging from 1.75-m (5.7-ft) to more than I1.0-m (36-ft) diam are available from several manufacturers; machines for mining applications generally have ranged between 1.75 and 6.0-m (5.7¬and 20.0-ft) diam. Tunneling machines usually operate at or near horizontal grades and are supported by a rail-haulage backup system. However, declines and inclines have been made using alternative backup equip¬ment. Either circular or shaped tunnels may be bored, depending upon the particular equipment used. Using tunneling machines to bore openings through rock offers several advantages over the use of the drilling and blasting technique. These advantages include a rapid advance, continuous operation, uniformity of the opening size, smooth walls, and safe working conditions. Compared to drilling and blasting, boring reduces the disturbance of adjacent rock, thereby reducing the ground-support requirements and the exposure to water inflows. The smooth walls provide natural arching and present minimum resistance to the flow of air. A tunneling machine should be considered a spe¬cialized mining tool that is well-suited to certain mining applications. Properly applied and sustained, a tunnel¬ing machine contributes to realizing substantial econo¬mies in the development of underground mines. All costs herein are expressed in 1977 US dollars, unless otherwise noted. The mention of trade names and manufacturers is not intended to imply an endorse¬ment of any specific equipment. TYPES OF TUNNELING MACHINES As listed in Table 1, five principal types of tunneling machines were being manufactured in 1977. Each type of machine offers a special solution to a specific mining problem. Although each of these types of machines is described herein, most of this chapter is devoted to the full-face rotary rock-tunneling machine that has demon¬strated the greatest production in both civil construction and mining applications. Full-Face Rotary Rock-Tunneling Machines As shown in Fig. 1, a full-face rotary rock-tunneling machine consists of a rotating head that is fitted with rock-cutting tools and is forced into the tunnel face. This head is supported by a bearing on a structural support member that, in turn, is held in place by a hydraulically positioned wall-gripping mechanism. Both torque and thrust are reacted from the cutterhead, through the structural support and gripper mechanism, and to the tunnel wall. The rock cuttings fall to the invert at the tunnel face, and they are removed by means of cutterhead buckets or scoops that transfer the cuttings to a conveyor belt. After advancing the cutter¬head through a preset boring stroke, the tunneling machine is advanced by hydraulically pulling the gripper mechanism in from the tunnel walls, stroking forward, and resetting the gripper to the new forward position on the wall; the machine then is set for the next advance stroke. Full-face machines cut only circular openings. Ma¬chines with diameters as small as 1.8 m (6 ft) and as large as 6.1 m (20 ft) have been used in mining applica¬tions, excavating rock with unconfined compressive strengths in excess of 276 000 kPa (40,000 psi). Although special machines have been designed to turn on a radius of 30 m (100 ft) in both horizontal and vertical directions, the conveyors and backup sup¬port equipment generally limit tunneling machines to turns with a radius of more than 60 m (200 ft). The full-face rotary rock-tunneling machines have provided the fastest and most reliable production of any mining method or machine. These machines can be equipped to place any of a variety of ground-support systems, and they can be provided as "open" machines, as shown in Fig. 1, or as "shielded" machines, as shown in Fig. 2. The open machines have the advantage of allowing the ground to be supported as near to the tunnel face as possible. The shielded machines have been designed principally to allow placing precast con¬crete segments. The shielded machines completely pro¬tect the equipment and personnel at the heading, but they are economically limited to permanent long-life haulageways where full lining is required. Partial-Face Rotary Rock-Tunneling Machines As shown in Fig. 3, the partial-face rotary tunneling machine is equipped with a cutterhead that rotates slowly at about 0.21 rad/s (2.0 rpm). The single-disk cutterhead produces the tunnel section shown in Fig. 4 by means of an oscillating motion. These machines work on the undercutting principle, and they are avail¬able in spring-line dimensional increments of 0.1 m (0.32 ft) from 1.5 to 2.1 m (5.0 to 7.0 ft); the vertical dimensions are 2.4 to 3.2 m (8.0 to 10.5 ft), respec¬tively. Depending upon the tunnel width, the cutterhead disk may be fitted with from 8 to 16 drag-bit cutters. Penetration is achieved by means of hydraulic rams that force the tungsten-carbide bits into the rock mass. The operation is cyclic, with cutting occurring only on the upstroke. On the downstroke, the rotating cutter¬head draws the fallen muck from the invert into the panzer-type conveyor for rearward transfer. This feature allows time to either reposition ore cars or to exit the tunnel with muck trains while the machine is cutting, as shown in Fig. 5. Boom-Type Rotary Excavators The boom-type rotary-excavator tunneling machines commonly are known as boom-type continuous miners

Jan 1, 1982

By Beth Robinson

This first of two articles on the Second Mexican Mining Rail Excursion describes some of the more casual experiences participants enjoyed on last fall’s rail tour of major Mexican mining operations. In this article, Beth Robinson, a graduate student at Colorado School of Mines, describes trip organizers, participants, and some of the unique events occurring during the two-week excursion. Next month, part two of the series will highlight various Mexican mines and plants.

Jan 1, 1982

By H. H. Richards, Ray L. Tobie, L. A. Thomas

GENERAL DESCRIPTION Since the beginning of operations, with the exception of a small tonnage mined by slushing, ore extraction has been by full gravity caving. Formerly, a checkerboard sequence of block undercutting was followed with the even-numbered blocks in one panel and the odd¬numbered blocks in the adjacent panel being mined. As these blocks were depleted, the intermediate or pillar blocks were mined (Fig. 1). Following this checker¬board, the mining sequence went through a number of changes, finally evolving into diagonal retreat panel cav¬ing by blocks (Fig. 2). The numbers in Fig. 2 indicate the sequence in which blocks were undercut. Gaps in the numbering sequence indicate undercutting on the level outside the illustrated area. Geology The ore body is a low-grade deposit of chalcopyrite mineralization disseminated throughout structurally weak, highly fractured, strongly altered granitic host rocks. It takes the shape of a gently dipping elliptical cylinder consisting of an ore shell of variable thickness surrounding an interior waste core. Major and minor axes of the mineralized cylinder are 1524 m (5000 ft) and 762 m (2500 ft), respectively, and length approximates 2438 m (8000 ft). Ore is sufficiently fractured to break readily into medium-coarse size. The igneous rock complex containing the ore body is covered by a wedge-shaped blanket of Tertiary con¬glomerate which was brought into place by faulting along the major regional structure of the San Manuel fault. Thickness of the conglomerate cap varies from only 9 m (30 ft) at the east end of the ore body to more than 610 m (2000 ft) at the west. Structurally, the con¬glomerate is much more competent than the igneous host rocks and, when caving, it breaks into massive chunks. Conglomerate boulders seen in drawpoints underground are very coarse. The total rock column over the initial mining area of the 1415 grizzly level was 354 m (1160 ft) of which 122 m (400 ft) was ore, 79 m (260 ft) was leached igneous capping, and 152 m (500 ft) was conglomerate above the San Manuel fault. Diamond Drilling: From 4572 to 7620 m (15,000 to 25,000 ft) are drilled annually from underground workings to delineate the ore body. MINE DEVELOPMENT Haulage Level In the south or main ore body (see Figs. 3-6), with the exception of the draw and transfer raises, all the extraction openings are concreted (Seaney and Tobie, 1965). The haulage panel drifts, which are 18 m (60 ft) below the grizzly drifts, are first driven with pre¬concrete ground support. The drift, which has an arched section, then is concreted using mobile collapsible steel tunnel forms. The haulage drifts leading from the pan¬els to the hoisting shafts are not concreted. After the panel drifts have been concreted, the raise stations from which transfer raises will be driven are constructed and the raise-station ore-drawing chute is installed. The chute is prefabricated of A-36 steel with undercut guillo¬tine gates made of abrasion-resistant 2.5-cm (1-in.) steel plate powered by 20-cm (8-in.) air cylinder installed on each side of the raise station. Transfer Raises The transfer raises are lined with 15 x 20-cm (6 x 8-in.) cribbing and are 1.22 m (4 ft) in the clear. Each cribbing is protected from wear by a high carbon steel angle which is nailed onto the cribbing. The transfer raises are driven from each side of the raise station on an angle of 1.1 rad (63°). Each raise con¬sists of a main and a backover branch. The transfer¬raise driving crew consists of two men working one shift only. Grizzly Drifts After the transfer raise reaches the grizzly level, the grizzly drift can be driven. The grizzly drifts are spaced at 10.6-m (35-ft) centers and are driven parallel to the long axis of the ore body (see Fig. 2). This drift is driven by a two-man crew working on one or more drifts at a time using feed-leg machines. The eight grizzlies in the 42.7-m (140-ft) long drift are spaced at 5.3-m

Jan 1, 1982