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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 David S. Bolin, Bruce Kennedy, Leigh A. Readdy, Graham Mathieson

INTRODUCTION The estimation of ore reserves is a process that begins with the earliest exploration stages on a property and continues throughout any subsequent evaluation and exploitation of the deposit. During exploration and preliminary evaluation, the results of these reserve estimates constitute the basic data for prefeasibility studies and economic analysis. The decision to continue exploration and development or to abandon a prospect is often based upon these studies. During the active life of a mine, reserve computations are continuously revised to assist in development planning, cost and efficiency analyses, quality control, and improvement of extraction and processing methods. Accurate reserve estimates are also required when financing a project, purchasing or selling a property, and for accounting purposes such as depletion and tax calculation. It is important to remember that the reliability of ore reserve estimates varies progressively through time as more and more information becomes available. The lowest order of reliability of estimation of reserves exists at the time of discovery. The maximum level of certainty concerning the ore reserves within a deposit is reached when the deposit is completely mined out. Between these two extremes are variable levels of certainty as to the tonnage and grade of the resource. This is particularly true of that portion of the mineral resource that constitutes the "minable reserve," as this portion is dependent upon economic as well as geological and technological factors. In the following discussion, several of the factors affecting ore reserve computation and some of the commonly used methods of calculation are presented. The first part of the discussion is confined to classical methods of hand calculation utilizing level maps and sections. The second part of the discussion presents an overview of somewhat more sophisticated methods of geostatistics which have been developed within recent Years. Geostatistical methods frequently provide the best ore re- serve estimates for many deposits. However, a digital computer is required for such geostatistically derived estimates due to the number and complexity of the cornputations involved. For many properties the classical nongeostatistical methods are adequate, particularly in the early exploration stages. In fact, it is the author's contention that even at those properties where the geostatistical approach is employed for reserve estimation, the geological and engineering staffs should prepare an estimate by traditional means. This provides an internal audit for the reserves and at the same time requires a continuing close appraisal of geologic problems that influence the presently known geostatistically estimated reserves and production at the mine.

Jan 1, 1982

By A. Fiszer

One of the major problems with deep Polish hard coal mines are deformations of close-to-wall galleries since they are necessary and must he kept open for emergency situations and are largely responsible for work disturbances. In order to ensure safe and effective working operations, it is necessary to often rebuild galleries which, apart from the inevitable stoppage of operations, is greatly disadvantageous for the operation of a mine. The largest problems with maintaining the galleries open occur in Lower Silesian Coal Field mines where sometimes it is necessary to reconstruct the galleries several times [(Fig. 1)]. In light of what has been said above, the aim of this article is to present the reader with methods and ways of determining deformations occurring in close-to-wall galleries. Squeezing or convergence of headings is only a portion of the entire process of ground movement caused by ground disturbance due to workings. The term "squeezing of headings" is to be understood as a shortening of the distance between a point set on the roof of a heading and the corresponding one on the floor. It is very important to know the rules governing the process of squeezing in a particular mining system since it affords possibilities to define changes of the process and its parameters at any moment for a given system. This is of utmost importance because convergence of a heading is one of the most significant manifestations of ground pressure which has already been proved by studies performed by Everling (1). On the basis of those studies it has been observed that there exists an almost linear relation between the rate of convergence and loading [(Fig. 2)]. This, in turn, means that the deformations of the already exploited portion of heading are proportional to the pressure changes.

Jan 1, 1982

By Richard L. Bullock

OPEN STOPING An open stope is an underground cavity from which the initial ore has been mined. Caving of the opening is prevented (at least temporarily) by support from the unmined ore or waste left in the stope, in the form of pillars, and the stope walls (also called ribs or abutments). In addition to this primary support system of open stoping, some secondary support may also be required using rockbolts, reinforcing rods, split pipes, or shotcrete to stabilize the rock surface immediately adjacent to the opening. The secondary reinforcement procedure does not preclude the method classified as open stoping. There are many forms of open-stope mining used to extract the initial material from a mine. Having once established that the mineral and waste rock are competent enough to use an open-stoping method, and assuming that the reserve is not classified as gassy, the Form which the method will take is primarily determined by the dip and thickness of the reserve. How these two factors affect the selection of the open-stope mining is discussed in a later chapter. At this point it will suffice to say that the classification of the open-stopes mining system which follows is based on whether dry' broken materials flows by gravity or whether it must be moved by nongravity methods where energy must be supplied to move the material. Room-and Pillar-Mining Room-and-pillar mining is an open-stoping method where mining progresses in a nearly horizontal or low angle direction by opening multiple stopes or rooms, leaving solid material to act as pillars to support the vertical load. Since the direction of excavation (angle of dip) is below that which would cause the dry material to flow by gravity to a drawpoint or gathering point, the material must be loaded in the room where it was extracted and transported to a point where it will flow, either by gravity or mechanical means, to a central gathering point to be taken out of the mine. This is an important aspect of room-and-pillar mining which differentiates the system from other open-stope mining methods which rely heavily upon gravity to transport ore from where it was broken to a lower elevation, usually through a drawpoint. There are many variations of the method which go by a number of names in local districts: breast stoping, breast-and-bench stoping, board-and-pillar, stall-and-pillar, and panel-and-pillar are all basically open-stope room-and-pillar mining. In some instances detailed stope planning is almost nonexistent; i.e., the operator simply follows the visual pay values, leaves pillars only where necessary, and tries to locate them in the zones of lower value. This method of mining is as old as the beginning of underground mining itself, dating back thousands of years. Early in the history of mining in this country, the term "gophering" was used to describe this method (Peek 1941). The term is appropriate, for it brings to mind the exact results of this type of system-a random and irregular room-and-pillar mine. In other instances where the mineral values are consistent both in physical dimensions and quality, the mine layout can be planned to the last detail, resulting in a uniform room-and-pillar mine. Coal, trona, gilsonite, potash, oil shale, salt, limestone, and sandstone mines can usually follow such a system. Today, most metal mines using a room-and-pillar operation try to mine as regular a pattern as possible but deviation in height, width, thickness, dip, and grade of the ore results in comparable deviation in the mine plan. Variations of the Room-and-pillar System It is necessary to briefly describe some of the many variations of the room-and-pillar system of mining, enabling the reader to fully explore the concepts and become familiar with the terminology used before going on to the details of mine design. Full-Face Slicing: If in the process of opening the rooms the total vertical extent of the mineral values of the particular seam or strata are extracted from the advance of one vertical face, the term used to describe this is full-face slicing. This face is also known as the "breast." There is no mineral of economic value intentionally left either in the floor or the roof (back) to be mined later. To be able to extract the full-face height in one pass, the mining equipment must obviously be designed to reach as high as the back. In an Appalachian coal mine, this may be all of 660.4 mm (26 in.); for a future oil shale property it might be 15.24 m (50 ft). Normally, however, in a majority of mines where the mining face gets over 6.09 to 6.7 m (20 to 22 ft), the tendency is to divide the face into more than one pass. Over this height, it becomes difficult to properly see and remove loose rock from the back with a hand "mining bar." Where the process of taking down loose rock has become mechanized, higher full-face mining can be safely practiced. Most eastern and midwestern coal seams and western uranium, trona, and potash seams in the United States are easily reached in a single face; many limestone, lead, and zinc mines must resort, at least in part, to "multiple-slicing" to remove all the minerals of value. Multiple Slicing (also known as multiple-pass mining): In many cases it is not practical to carry the full vertical height of the mining horizon as a full face. The face is divided into parts known as the breast, bench, and/or brow. Ideally, if the operator knows the vertical extent of the mineralized zone, he will drill and blast the first pass at the top of the zone, thereby

Jan 1, 1982

By S. McMurray

INTRODUCTION Towards the end of the 1890s the attention of a prospector by the name of Thomas Cullinan was drawn to the occurrence of alluvial diamonds east of Pretoria. Persistent prospecting work by Cullinan led to the loca¬tion of the source of these diamonds-a kimberlite pipe on the Elandsfontein farm. Cullinan eventually acquired the property in 1902, and during an extensive evaluation program outlined the largest diamondiferous pipe in South Africa on which the Premier mine was established. Worldwide attention was drawn to this mine in 1905 when the Cullinan diamond, which weighed 3106 carats and is by far the largest gem diamond ever discovered, was recovered. Premier mine operated with varying success until 1932 when the worldwide depression forced it to close. In 1945 the mine was reopened and since then has been a major producer of gem and industrial diamonds. During the early 1950s exploratory drilling revealed that below the 370-m level the pipe had been completely cut off by a younger gabbro sill. This sill has created major mining and metallurgical problems but the existence of vast reserves of high grade ore below it justi¬fied the establishment of virtually a new mine below the sill. Detailed planning of below-the-sill mining is nearing completion and the development of the first mining block (the L1 block) is well advanced. A full discus¬sion of this block follows. GEOLOGY OF PREMIER MINE Pipe Morphology and Country-Rock Geology As can be seen in Fig. 1, Premier mine is an elon¬gated oval shape with a long axis of 900 m on the surface and a short axis of 450 m. The pipe has a surface area of 32 hectares which decreases progres¬sively with increasing depth, so that 500 m below the surface the area is reduced to 22 hectares. The contact between the kimberlite and the surrounding country rock is sharp, with an average angle of dip of 1.48 rad (85°). From the surface down to approximately 350 m, the country rock is a felsite which grades downwards into a norite. In general both rock types are hard and massive, the felsite in particular being poorly jointed. This characteristic combined with other factors results in an extremely stable open pit, and near-vertical side¬walls are currently being maintained. The norite in general is competent but is heavily jointed in some areas resulting in very blocky ground which creates localized tunnel support problems. In the southeastern area below the sill a zone of highly altered and very unstable norite, which has also created localized sup¬port and development difficulties, lies adjacent to the pipe contact. Kimberlite Geology The Premier pipe is a complex multiple intrusion which contains in the region of 15 separate types of kimberlite, most of which are volumetrically insignifi¬cant. On a simplified basis three major types can be identified which correspond to three separate phases of intrusion. They are: Brown Kimberlite: This represents the first phase of intrusion and now occupies the eastern part of the pipe. In plan the brown kimberlite has a crescentric shape resulting from the intrusive relationship with the younger gray kimberlite which lies to the west. The brown kimberlite increases in relative area with increas¬ing depth which is a favorable factor for future mining operations since this kimberlite is the richest area of the mine, having an average grade of 70 carats/ 100 t. Below the sill this rock disintegrates extremely rapidly when exposed to air or water, to form a fine gravel-like material. This characteristic is due to a high montmorilonite clay content. As is common with clay minerals of this type, water can be absorbed into the crystal lattice with a resultant increase in volume which causes the physical decomposition of the rock. The presence of this decomposing kimberlite was a major factor governing the choice of a mining method, as well as development and tunnel support techniques. Gray Kimberlite: This is the most abundant type and represents the second major phase of intrusion. The diamond content is variable depending on the amount of waste rock dilution, and in general the rock is stable and non-decomposing. At the time of formation of Premier, the surface rock type was Waterberg quartzite, a rock formation which has subsequently been removed from this area by erosion. The explosive extrusion of the gray kimberlite caused extensive brecciation of this rock and while the kimberlite was in a mobile state huge masses of quartz¬ite, amounting to tens of millions of tons, slumped into the pipe. This quartzite became concentrated in the central area of the mine and has been a major mining problem throughout the history of Premier since, due to the irregular distribution of the quartzite, effective selective mining has never been possible. The quartzite is therefore extracted with the ore and has a significant dilution effect on the overall grade. Black Kimberlite: In the western part of the pipe a circular, pluglike body of kimberlite is intruded into the gray kimberlite. This plug consists mainly of a hard, dark-colored rock known as black kimberlite, but also contains a number of minor kimberlite variet.ies as well as nonkimberlite carbonate dikes. The Gabbro Sill The sill averages 75 m in thickness and within the confines of the pipe consists of 52 million t of rock. On average the major dip is 0.34 rad (20°) to the northeast which is across the short axis of the pipe, the range of dip being from near horizontal up to 0.52 rad (30°). Due to the pronounced dip, the sill is encount¬ered over a vertical distance of 175 m with the highest upper contacts on the 355-m horizon and the lowest bottom contact on the 530-m horizon.

Jan 1, 1982

By Frank G. Horino, V. E. Hooker

INTRODUCTION The determination and use of mechanical properties of rock in engineering and rock mechanics are rapidly developing. Many of these properties are determined on intact rock specimens; thus, their application and repre¬sentation of rock mass properties may be limited. How¬ever, relative information often provides useful guidance in the solution to mine design and stability problems. Summarized in this chapter are some of the stan¬dardized techniques and procedures currently used to obtain these mechanical properties. Typical applications of the use of these properties are also presented. Stan¬dardized techniques include those advanced by the American Society of Testing and Materials (ASTM), International Society of Rock Mechanics (ISRM), US Bureau of Mines (USBM), Canadian Dept. of Mines and Technical Surveys, South African Institute of Min¬ing and Metallurgy, and other individual investigators. Information on the mechanical properties of rock and the behavior of the rock under a given system of stresses represents a necessary part of the information for rational engineering design for any given mining op¬eration. The mining method, the type and extent of sup¬port, the extraction ratio, the overall dimensions of the mine, and the orientation of the rooms and pillars are all decisions that are influenced by the mechanical prop¬erties of the ore, roof, and floor material under various stress systems and the magnitude and direction of the in situ stresses (Hooker, Bickel, and Aggson, 1972). Initial mechanical property information regarding a structure or mine property is generally obtained by two basic techniques: (1) static and dynamic property tests are conducted on intact and fractured rock specimens of exploratory drill core, and (2) dynamic properties are obtained by borehole logging techniques. When mining access becomes available, and as the mining horizon is expanded, additional information can be ob¬tained to verify preliminary mine design values. This chapter presents some of the standardized tech¬niques and equipment currently used in obtaining me¬chanical property data in the laboratory. The properties considered are: (1) uniaxial compressive strength of intact rock core specimens, (2) uniaxial compressive strength of rock cores containing planes of weakness, (3) triaxial compressive strength of intact rock core specimens, (4) triaxial compressive strength of cores with a plane of weakness, (5) Young's modulus, (6) Poisson's ratio, (7) density or apparent specific gravity, (8) modulus of rupture, (9) indirect tensile strength, and (10) creep characteristics. Where possible, an at¬tempt will be made to evaluate each property measure¬ment in relation to the problems of rock mechanics and application of results. TEST SPECIMENS The selection and care of drill core for laboratory testing require some consideration. It is recognized that laboratory-determined properties are not necessarily rep¬resentative of an in situ rock mass property. However, relative information between beds or zones of interest is still valuable information in selecting mining horizons and preliminary design criteria. To provide statistical data the number of drill core samples selected to repre¬sent each of the areas of interest should be from a mini¬mum of three to a maximum of ten test specimens. A judgment must also be made on site as to whether the recovered drill core should be wrapped and sealed in plastic to preserve moisture. On the one hand investiga¬tions of air-dried and saturated specimens have shown that moisture significantly affects the elastic properties and strengths of many rock materials (Obert, Windes, and Duvall, 1946; Colback and Wiid, 1965); on the other hand it is apparent that most core drilling is done with water which may saturate the specimen to a greater extent than in the in-situ condition. Whether or not the decision is made to retain the moisture, the core should be delivered to the laboratory as soon as possible after recovery for subsequent specimen preparation and testing. Specifications Shape: The shape of the specimens influences lab¬oratory testing in two ways: (1) time and cost of sam¬ple preparation and (2) strength of the material. Cy¬lindrical specimens of drill core are by far the least time-consuming to prepare for static or dynamic labora¬tory testing. In addition, the cylindrical shape lends it¬self to a more uniform stress distribution throughout the sample than other shapes, such as rectangles and hexa¬gons. The compressive strengths of various shapes have been studied (Grosvenor, 1963, and Price, 1960), and results indicate that the cylindrical specimens usually provide the highest strength for a given height-diameter ratio. However, reduction in strength from a cylindrical shape to a rectangular in situ pillar is not regarded as significant in relation to other considerations such as planes of weakness in a pillar or safety factors in the design process. Length-Diameter Ratio: The length-to-diameter ra¬tio, LID, has a significant effect on the compressive strength. Various recommendations have been made to use standard LID ratios ranging from 2 to 2.5 to 3 (ASTM, 1975c; ISRM, 1972). However, past work by others such as Obert, Windes, and Duvall (1946) has shown that excellent results can be obtained using LID ratios from 2 > (LID) > >/s. In selecting an LID ratio for testing, one should keep in mind the amount of material available for testing. In many instances, this may be limited. Thus, a shorter specimen such as 1: 1 LID may be necessary to provide enough test data for statistical analysis of results. Sec¬ond, it may be desirable to obtain elastic constants dur¬ing the test. This generally requires instrumentation such as linear variable differential transformers (LVDTs) or strain gages near the center of the specimen. In this case, an LID of 2.5 or 3 is desirable so that the instru

Jan 1, 1982

By Rex C. Bryan

An application of geostatistics to financial decision analysis is the linking of reserve parameters estimation precisions to a mine economics computer program calculating profitability. The estimates of economic reserve parameters are obtained by the kriging method, while precision are calculated by a technique more akin to the polygon method. To indicate the effect of the variables probability distribution on a distribution of profitability, their values are sampled by a Monte Carlo technique and entered into a mine economics modelling program developed by T. R. Ellis. The distribution of profitability calculations are in essence, a superior method of sensitivity analysis. The method is applied to a roll-front uranium deposit in the Powder River Basin, Wyoming.

Jan 1, 1982

By Robert L. Sundeen, Richard V. Wenberg

INTRODUCTION Early rock-bolting practice usually consisted of ran¬domly installing bolts, using hand-held drilling equip¬ment such as air-leg mounted sinkers or stopers. As bolting technology evolved, the practice of installing uniform bolt patterns was developed, utilizing random bolt installation as required within the uniform pattern. The uniform rock-bolting technique, when used as the principal method of rock support, requires drills and jumbos designed specifically for rock bolting. This equipment is similar to conventional face drilling machines, except that the jumbos can orient the drills to obtain any necessary bolting angle; often, this means bolts may be installed at or near perpendicular to the centerline of the drift or stope opening. Fig. I shows a typical rubber-tired rock-bolting jumbo of the type commonly used in drifts and room-and-pillar mining. Modern rock-bolting jumbos are capable of drilling the required holes, inserting rockbolts, and torquing them to proper levels. Jumbos also are available for use with resin-anchored bolts. Probably the most unusual of this type utilizes a rotating bolt turret that carries the rock drill, resin charging unit, and the bolt in a single mechanism. Once the turret is loaded with the resin and bolt, further manual effort is not re¬quired. Drilling and bolt placement are accomplished by the operator from a station remote from the bolt site. Fig. 2 shows this turret device. Roof-bolting jumbos are available mounted on rail, crawler, or rubber-tired chassis. Various types of rock drills are used. Depending upon the type of rock, these may include rotary, percussion, or rotary-percussion drills. Fig. 3 shows a typical rail-mounted jumbo. CONDITIONS INFLUENCING APPLICATION General Rock Conditions Rock-bolting generally is applicable in moderately hard to hard rock. No particular type of rock is ex-

Jan 1, 1982

By Charles J. Hass

INTRODUCTION Deformation measurements are some of the simplest and most useful measurements that can be made by the mining/rock mechanics engineer in assessing the sta¬bility of openings in rock. A number of tools and tech¬niques are presently available. This chapter will de¬scribe the most common instruments, their application, and present the equations needed for data reduction. MECHANICAL STRAIN GAGE Description of Instrument A mechanical strain gage of the Whittemore type shown in Fig. I is a useful instrument for measuring deformations of pillars, sidewalls, and roofs of under¬ground mines. Measurement gage lengths typically range from 51 to 254 mm (2 to 10 in.), and the dial indicator reads to 0.0025 mm (0.0001 in.). The instru¬ment measures the deformation between precision gage points which are anchored in the rock. The anchor points are typically grouted in small holes drilled 12.5 mm (1/2 in.) or so deep into the rock. Strain a is calculated from the measured deformation AL with the equation [ ] where L is gage length. A necessary accessory for the instrument is a master bar normally made from invar steel having a low co¬efficient of thermal expansion. This bar allows one to check the instrument for changes in the mechanical zero point and to correct for temperature effects on the measuring instrument. The instrument is available from Soiltest, Inc. for a cost of approximately $340; the cost of the master bar is approximately $55; and nonreusable brass inserts which are anchored in the rock are nominally priced at $0.27 each. All prices are for 1978. Applications Two classes of applications may be considered, one where only one measurement direction is involved and the second where three measurement directions are em¬ployed. Applications where only one measurement di¬rection is necessary include: monitoring crack or frac¬ture separation or shear movement with anchor points on opposite sides of the discontinuity; measurement of vertical strain in a pillar or sidewall; and measurement of steel support deformations with anchors in the steel members. The more general application involves a strain ro¬sette consisting of three measurement directions which for convenience are oriented 1.04 rad (60°) apart to form an equilateral triangle as shown in Fig. 2. Strains in the three measurement directions [ ]

Jan 1, 1982

By William F. Eichfeld, Robert J. Wenzel

INTRODUCTION The phrase "subsidence control planning" can refer to the use of specific mining methods and-specialized techniques to control subsidence events or it can refer to planning for subsidence by predicting its magnitude and timing and further planning mitigation strategies. In the first case the objective is actually subsidence prevention while in the second it is a design procedure that seeks to maximize the use of both surface and subsurface resources. Neither technology is well developed at this point. The purpose of this presentation is to briefly outline the Department of Energy's Subsidence Program. This program is, in general, bases on the premise that subsidence control technology, by any definition, can be accomplished only through a thorough and rigorous understanding of subsidence phenomena. This understanding must come from many well planned and executed studies of stress redistribution and strata deformation as they occur during mining activities. There is presently considerable incentive to improve our subsidence engineering capabilities due to the desire for increased coal utilization with acceptable environmental impact. Considerable research in this area is being sponsored by federal and state governmental agencies, universities, and the coal industry. Federal sponsorship has been borne by the Department of Energy, the Bureau of Mines, the Office of Surface Mining and the U.S. Geological Survey for the most part. These agencies have carried out basic research in their laboratories and have contracted with other organizations for additional work. At this time most of the available resources are directed toward contract research, especially in DOE. It is the DOE research plan which will be addressed here.

Jan 1, 1982

By Paul T. Allsman

The guiding principle in developing this plan resulted from evaluating potential future markets for oil, the quantity of mineable oil shale and the optimum rate of production comensurate with market demand on the available oil from shale. In addition, the rate of production may depend upon a solution to the problems of scarcity of resources other than oil shale. The rate of production finally will tend to establish the choice of technologies with these limits. The principle components discussed will be: 1) The place of oil shale among the various energy sources and its relative adaptability for use in the area of major consumptive demand. 2) A recalculation of the reserves and their classification on the basis of quantity, grade - gallons per ton, - and physical characteristics principally thickness, overburden and depth.

Jan 1, 1982

By S. S. Peng

In dealing with the mining engineering problems, the traditional approaches of laboratory experimental simulations and/or analytical solutions are not very effective. Mainly because the field problems are too complex to be solved with a few simplified assumptions. Conversely it is also true that the experimental and analytical/ numerical models currently available are far from being able to model the actual cases. Under such conditions, the most effective mean of research into mining engineering problems is to combine the mine (field) investigations with laboratory and / or analytical modelings. It is with this concept in mind this research was initiated in 1973. 'The field investigations which began in the fall of 1978 consisted of mainly visiting the mines where surface subsidence monitorings had been or were being monitored and/or surface structural changes had occurred. From such visits, subsidence data were collected. A quick analysis of those data provided preliminary understandings of the characteristics of the surface subsidence parameters in the regions (1-5). Using those date, models of various types are being developed for subsidence prediction / control. So far, a total of 44 longwall subsidence profiles in 22 longwall panels and 10 room-and-pillar profiles in 5 room and-pillar sections have been collected. This paper summarizes the combined results published previously (1-5) and present the preliminary results of the predictive models developed.

Jan 1, 1982

By J. van Oldenborgh

The world market for bunkers represents some 140 MT (1980) of marine fuel oil and distillates. This is equivalent to 23CMT of coal, a potential market for steam coal larger than the total seaborne trade of steam and metallurgical coal today, and an obvious area which could contribute savings in world oil consumption. Oil bunkers are lifted fairly evenly throughout the world, the main supplying regions being Western Europe, the Middle East, North and Central America and Japan (Figure 1). Not surprisingly, the main customers are the tankers representing about 50 percent of the world market - (95 percent of the Middle East Market) - dry bulkers represent some 30 percent and general cargo ships the rest.

Jan 1, 1982

By Alvaro Lucio

The continuous increase of the fuel oil price compelled CV3D to look for an alternative fuel. Charcoal was selected as the alternative energy source. This paper justifies its selection, describes charcoal properties, the methods for using charcoal in the pelletizing and shows the results of the in-plant tests.

Jan 1, 1982

By Dr. Peter J. Conroy

The current study is part of the U.S. Department of Energy's (DOE) on-going subsidence research program. The long-term objective of the DOE program is to develop analytical methods of subsidence prediction applicable over a broad range of geologic and mining conditions encountered in the United States. The DOE currently has several contracts in progress to develop a sound data base that can be used in the development of predictive models. Dames & Moore is currently under contract to the DOE to collect and analyze detailed subsidence and hydrologic data from a coal mine employing longwall mining methods in the Eastern Coal Province. Work under Contract No. DE-AC22-80PC 30335 began in October of 1980 and will continue through September of 1984. This paper summarizes the work that has been completed through September of 1981 and presents the results that have been achieved during this period. Technical difficulties encountered during both the installation and monitoring phases are discussed, and recommendations for future work are presented.

Jan 1, 1982

By Allister McLeod

Coal is expected to play a major role in energy supply in the next 20 years. Oil is a convenient and still reasonably abundant fuel. However, its major sources of supply are regions where there is nervousness about political stability. Uranium is a more efficient fuel - but there is a continuing community apprehension regarding its use. What coal may lack in efficiency, it makes up in long term availability and competitive overall cost. Moreover, diversity of source ensures stable long-term supply. In this way, a major role for coal in the world energy scene is assured. Cyclical considerations aside, demand for coal will increase significantly and it will meet a substantial share of new energy demand, replacing oil and gas in major industrial and electric utility markets. By around the year 2000, coal's share of world energy could reach 30% and rival oil as the largest single source of energy.

Jan 1, 1982

By Evans U. Jegbefume

A plane strain two-dimensional finite element model employing the method of negative reaction loading has been used to model roof collapse and subsidence resulting from the underground gasification of coal in regions of soft overburden such as that found in Texas. The model included dry zones of rock around the cavity. Roof collapse was included by considering the final state of stress in the elements and removing those that had undergone tensile failure. The two-dimensional plane strain model was extended to include thermal loading and thermal softening of the roof rock. Roof collapse initiation and growth of the cavity were modelled by examining the portion of the rock immediately above the cavity under thermal and hydrostatic load. The model was extended to include linear creep using the method of incremental strains. The overburden was modelled as a generalized Kelvin material.

Jan 1, 1982

By Donald L. Bott

A computer model called the Quality Compliance Model was developed to optimize the mine planning and sequencing in area coal mining with respect to the limit on pounds of sulfur dioxide per million BTU imposed by regulations. The objective of the model is to maximize the amount of high sulfur coal mined from various benches while meeting the specified compl'ance standards. Given multiple seams and multiple benching operations, the mining sequence from each pit is simulated. The rate of mining is controlled such that the in-pit blending will maximize the usage of high sulfur coal and meet compliance standards. The compliant coal can be stockpiled for usage at a later date, or it can be disassociated. This calculation can be based on any given rate of production, i.e., from as small a unit as a unit train to as large as monthly or yearly production.

Jan 1, 1982

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By Ronald E. Yarbrough

As a result of concern for the citizens of the State of Illinois who owned structures above underground mine workings, the Illinois State Legislature passed House Bill 158 in 1978. This bill amends the Illinois lnsurance Code and creates the Mine Subsidence Reinsurance Fund which is an insurance pool that is to receive a state appropriation of $500,000 as "start-up" money. One hundred thousand dollars has been paid to the fund as of October, 1982. The Fund is to return this appropriation to the State and is to remain self-supporting from premiums. The Fund utilized a pre-existing institution, which also was an insurance pool, the "Fair Plan Facility" and its governing Board, consisting primarily of insurance executives, to administer the program. The Mine Subsidence lnsurance Act went into effect October 1, 1979, and the bill provides that property owners in the State of Illinois may purchase up to $50,000 of mine subsidence insurance which is attached to their fire insurance policy. The deductible is normally 2%, but cannot be less than $250 or more than $500. The annual rate structure for residential structures in 1981 for $50,000 coverage is $10 (frame), and $12 (masonry). The rate structure for commercial buildings is based on the value of the building and the rates for $50,000 maximum coverage is adjusted upward. The insurance was mandated by the State Department of lnsurance and rolled onto the insurance policies of the homeowners in 34 counties which had been recognized as having sufficient exposure to subsidence damage to be covered by the Act. One percent or more of the land surface is undermined in each of the 34 counties. In the other counties undermined in the State of Illinois, insurance is available to the homeowner upon request to the private insurance carrier.

Jan 1, 1982