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By Ivor Hawkes, Malcolm Mellor

INTRODUCTION There are currently around fifteen companies in the United States and Europe manufacturing hard rock tunneling machines. In recent years upwards of 200 hard rock tunnels of lengths up to 12 miles and diameters to the order of 30 ft have and are being successfully driven by over 100 different machines. It has been predicted that within the next decade over 2,000 miles of tunnels will be driven for transportation purposes in the United States alone (OECD Advisory Conference on Tunneling, Washington, 1970) and it is expected that eventually most tunnels including mine drivages will be excavated by machines. All tunneling machines have certain common basic functions; thrust is applied to drive or hold the rock cutting tools into the face, and torque is applied to rotate the cutting tools over the face so that they can continuously break out rock. The speed of head rotation, together with the torque requirement, determines the cutter head power. (The power required to push the head forward is usually negligible.) There are many factors which determine how successfully a given tunneling machine will perform in a given rock but one of the most important is effective utilization of cutter head power. This paper is a review of the specifications and performance of hard rock tunneling machines, obtained from manufacturers' catalogs, published literature, talks with manufacturers and visits to operating machines throughout the United States and Europe.

Jan 1, 1997

By Siegmund Babendererde

Introduction Lyon, with 1.1 million inhabitants, is the third largest city in France after Paris and Marseille. It is a lively industrial centre. Lyon lies at the intersection of many transport routes and at the fluence of Rhone and Saone. Like all large cities, Lyon is attempting to alleviate inner city traffic problems by constructing a metro. Metro construction in Lyon began with tunnels built by cut and cover methods along the streets. Deep diaphram walls, which were back anchored and toed into imperveous strata, formed the excavation walls. It is a very economic and simple method. However during the construction period, it has the disavantage of strangling urban life for several years. It is therefore increasingly being replaced by subsurface tunnelling methods. As it was decided to build the metro tunnel between the Guillotiere and St. Jean stations, and thus cross under the Rhone and Saone, the client, Semaly, put the contract out for a competitive tender in which the general parameters were defined. Line, level and construction method were left to the contractor. There was an extensive examination of the tenders, during which reference projects from competitors in Japan were visited, as well as in Germany. Finally the consortium of Dragages, Campenon Bernard, Colas and Hochtief was awarded the contract to drive this tunnel section with an hydro-shield and an extruded steel fibre concrete lining, a Hochtief design.

Jan 1, 1987

By David T. Swanson

The term "pressure tunnel" in the broad sense includes all hollow spaces, aligned along an axis and surrounded by rock, which are suitable for pressurization. In particular, this category includes: • horizontal or slightly inclined water pressure tunnels; • vertical or inclined pressure shafts; • pressure chambers or shafts of surge chambers; and, • compressed air storage tunnels. The main problems in the design and construction of pressure tunnels result less from their length or size than from the requirement for watertightness and, dependent upon this, reliability in operation and economy in construction. Where the lack of sufficient overburden does not permit the rock to accept the internal pressure in the tunnel, or where this pressure is so high that the watertightness is in doubt although the stability of the tunnel shell is not in question, a pressure tunnel must nevertheless be reinforced. Surge chambers, pressure tunnels and tailraces which are subjected to relatively low internal pressure, usually not much in excess of 1.5 MN/m2, are eminently suitable for construction in post-tensioned concrete. Pressure shafts, which as a consequence of the high water head are subject to a considerable maximum service pressure, are usually lined with a steel tube, a considerable proportion of the internal pressure being transmitted to the rock by backing concrete. TYPES OF LINING FOR PRESSURE TUNNELS In general, the type of lining will depend upon the nature of the rock. As a result, pressure tunnels usually are not uniform in con-

Jan 1, 1981

By A. A. Mathews

INTRODUCTION When Palmer King asked me to prepare a paper for this conference, it seemed quite appropriate, since I am Chairman of Subcommittee No. 4 of the U. S. National Committee on Tunneling Technology and we are presently making a detailed study of contracting practices. Before the end of this year, we expect to publish our report containing guidelines aimed at improving our contracting practices and minimizing the problems which develop therefrom. The American construction and mining industries today are glowing examples of the ingenuity of man pitting his wits against the unknown. Without the promise of rewards commensurate with the perils faced and the effort expended, it is doubtful that these industries could have reached their present state of advancement. Nevertheless, these industries are still primitive compared to many others such as aerospace and electronics. That our present contracting practices leave something to be desired is evidenced by the fact that many of our underground projects wind up with tremendous increases in estimated cost, financial disaster, and costly dispute and litigation. The situation is aggravated by the current economic uncertainty, energy crisis, and shortage of materials. At the same time, if our industries are to develop their maximum technologic potential, we must employ contracting practices which will encourage that development. Most of the contractual problems related to underground construction are associated with risk and its management. Unless all of the ramifications of this subject are understood, it will be difficult

Jan 1, 1974

By D. D. Snyder, J. D. Bode, F. T. King

INTRODUCTION Only in recent years have the environmental hazards associated with mining and tunneling operations been fully acknowledged and determined remedial action undertaken. The nature of these excavation operations leads to a wide diversity of hazards spanning mechanical, electrical, chemical, environmental, and physiological. The most insidious of these are unquestionably the environmental ones which have the greatest long-term effect and are the most difficult to pinpoint and monitor. It is the atmospheric contaminants that lead to a significantly greater illness and mortality rate among miners. Dust is the most prevalent, hence most important of the atmospheric contaminants. Respirable particles lead to diseases characteristic of many mining operations: coal workers pneumoconiosis, asbestiosis, silicosis, talcosis, anthrocosis, etc. But toxic and suffocating gases can be equally insidious and the most innocuous pollutant, methane, leads to the most imminently disastrous results of all - explosion and fire. This means that truly effective monitoring of the environmental hazards in tunneling cannot be limited to a few simple standard check measurements. It must include the most precise measurements we can make for specific, documented hazards and even for factors, such as noise and heat stress, not immediately obvious as hazards. The Federal Metal and Nonmetallic Mine Safety Act and more recent Occupational Safety and Health Administration standards are clearly designed to match such effective monitoring. We must, therefore, consider not only presently available monitoring instruments and those currently under develop-

Jan 1, 1997

By Kerry Spickelmier

A traveler driving across desolate central Nevada to the Round Mountain property can easily identify with turn- of-the-century prospectors seeking a fortune, or at least an existence, from the high desert. But now a shadow passing overhead is as likely a sophisticated fighter aircraft from Stillwater Naval Air Station as a circling bird, and the famous trails west show only as historical markers along the highway. Even the approximately 400- km (250-mile) trip southeast from Reno or northwest from Las Vegas takes hours now, not days. And the camps of Round Mountain’s early prospectors have been replaced by one of the world’s largest gold heap-leaching operations. While Round Mountain’s uniqueness begins with its scale, it doesn’t end there. Consider the mine’s low cutoff grade of 0.2 g/t (0.006 oz/st) gold, made possible by a new dedicated (not reusable) leach pad. Also, the operation’s degree of automation means that ore dropped in the south crusher system is not manually handled until it is stripped from the pads. There is the bonus of a geothermal field that is tapped to heat the leach solutions in the winter. And, of course, there was that “unexpected” hitch expected in all mining operations. For Round Mountain, it meant buying a gravity plant to process the gold nuggets discovered last summer in a fault gouge.

Jan 1, 1993

By Usko Anttikoski

This article deals with the extensive geotechnical soil and bedrock mapping system maintained in the Geotechnical Office of the City of Helsinki since 1955.Soil maps have been prepared to scales of 1:500,1:2000 and 1:10 000, and most recently in 1978 a new type of bedrock map was printed to a scale of 1:10 000, which covers the whole city area and particularly serves rock tunnel construction. A map showing the types of foundation building in the central city area, which will also serve underground construction activities, is currently being prepared. Since the geotechnical mapping system has been well received in Helsinki and the drawing of similar maps has also been undertaken in other Finnish towns, this mapping method is believed to be of interest to foundation engineers and tunnel constructors elsewhere too. THE GEOTECHNICAL CONDITIONS IN HELSINKI The geological structure of Helsinki, and of all Finland, presents two types clearly differing in character and age; bedrock and soil. The solid bedrock is Precambrian, while the soil dates from the Pleistocene Age and has mainly been deposited during and after the Ice Age. Thus the history of Helsinki only displays very old and very young formations, but nothing in between. The ground level in Helsinki is +62 at its highest and the sea bottom close to shore is -15 at its deepest. The deepest rockhead has been found to lie at the -70 level.

Jan 1, 1979

By Max John

INTRODUCTION One of the main advantages of the New Austrian Tunneling Method (NATM) is its adaptability to varying rock conditions. This method is therefore successfully applied when complex geologic features are expected which cause uncertainties in the prognosis of the rock mass behaviour, as was the care at the Arlberg Expressway Tunnel. Furthermore, by employing the NATM problems are avoided when encountering varying geologic conditions or even extension fault zones along the tunnel, as was the case at the Dalaaser Tunnel and Perjen Tunnel. The NATM can also be used for rock and soil. The following examples will show how various rock conditions can be mastered. In accordance with the NATM all tunnels which will be discussed are road tunnels having a cross section area between 70 and 100 m2. The construction consists of a primary support, which ensures stress redistribution during excavation (Figure 1). The primary support is reinforced shotcrete strengthened by steel arches, when necessary, and rock bolts. When encountering rock pressure, swelling rock or soil, the primary support is suplemented by an invert arch. The purpose of the secondary lining is to increase the safety of the tunnel with regard to long-term stress redistribution processes or changes in rock mass properties as well as to protect the insulation. The secondary lining consists of cast concrete without reinforcement and is generally 0.25 or 0.30 m thick.

Jan 1, 1981

By Robert W. Miller

There are many problems faced by ventilation engineers in deep underground mining operations, not the least of which is controlling miner exposure to radon gas and its daughter products. Radon gas is commonly found in uranium mining operations, but may also be present in other deep metal mines. For example, tin mines in England, iron ore mines in Sweden, gold mines in South Africa, and molybdenum mines in the U. S. have potential radon exposures. This is because uranium and accompanying radium ore are ubiquitous to the earth's crust albeit at low levels. The fact that the activity represented by one WL can be caused by a relatively low concentration of radon gas increases the difficulty of control. Since the source of the radon gas is usually widespread throughout a mine, local exhaust ventilation is not a viable control schema. The technique used to control exposure is then dilution ventilation and, in fact, huge amounts of air must be moved in order to reduce potential exposures to an acceptable level. An interesting comparison can be made of ventilation rates in different types of mines. It is estimated in modern coal mines, which are generally acknowledged to have high rates of ventilation, that about eleven tons of air are moved for each ton of ore mined. A typical operating uranium mine may have ventilation flows of 14-15 tons per ton of ore mined. This provides an idea of the scope and importance of ventilation in modern mining operations where radon is a hazard. Further pressure is put on ventilation engineers by the steady downward trend in exposure limits set by national and international standard setting agencies. Much of this tendency toward lowered standards is based upon longitudinal mortality studies of miner populations. Another important factor is the limited number of experienced miners available in the labor pool. For optimum production, it is important to have as many experienced miners underground in each shift as possible. However, the average daily exposure in a U. S. mine must be less than .3 WL to permit the miner to work underground for a full year. The ventilation system then must provide enough uncontaminated air to maintain the WL below the .3 TTL level to maximize production efficiency and minimize personnel turnover and the problems associated with it. Ultimately, the goal of the ventilation engineer and health physicist is to protect the working miner from harmful exposures based upon currently acceptable standards. U. S. Federal regulations require that in uranium mines all active work sites must be monitored every two weeks if they measure above .1 WL. Areas that have .3 WL ratios or higher must be monitored on a weekly basis until five consecutive weekly samples show the level has dropped below .3 WL. Also, exposure records must be kept for all individuals exposed to levels exceeding .3 WL. These requirements provide a strong economic incentive to have a ventilation system that minimizes exposure of any personnel. A good ventilation system requires careful planning, operation and backup in order to fulfill its mission of providing adequate clean air. Its proper operation also requires coordination with production personnel so it can be adapted as new areas in the mine open up and old areas are sealed off. The ultimate indicator of ventilation efficiency to control radon daughter exposure is, of course, monitoring working levels. Historically, this has been done using the Kusnetz, Tsivoglou, and Rolle's methods, among others. These methods all require cumbersome equipment and tedious calculations to obtain the measurements that results in WL. More important, however, they require a significant time lag between sampling and counting, typically 40-90 minutes. This time lag is, in fact, what can cause significant economic losses due to unnecessary downtime as well as high WL exposures. In a typical mining situation, a sampling technician using the Kusnetz method takes a sample, moves to the next location and takes another sample and so on. Forty to ninety minutes after the first sample, the technician will stop, run the activity count on the filter and calculate the WL. The technician may be one-half mile away or several levels removed from where the first sample was taken when it is counted. If the WL ratio is high the technician must then backtrack to the sample position. There are then two options. If the sample area is a working stage, it can be shut down or a second sample can be taken. If the first alternative is chosen; i.e., shutdown and correction of the ventilation, then another sample must be taken, followed by a forty minute wait for results. If the ventilation adjustment didn't correct the problem, then the whole process must be repeated with a minimum of forty-five minutes per sample cycle when using the Kusnetz method. It has been estimated from operating uranium mines that the cost per hour for downtime on a production slope is about $1,50O/hour. The time lag between sampling and resultant data can be very costly. If the second alternative is chosen to verify the first reading, the miners may be unnecessarily exposed to high levels while waiting for the result. Clearly, such a sampling system can be markedly improved by eliminating the excessive time lag between sampling and analysis.

Jan 1, 1981

By Wallace Jarman

Introduction As with many gravity concentration processes, pneumatic concen¬tration traces its origin in antiquity. The use of winnowing to separate chaff from grain has long been known, and such procedures were undoubtedly practiced in ancient civilizations for the concentration of ores and the separation of slag from metal whenever specific gravity, size, and shape differences were favorable. Taggart142 has reported on the use of dry rockers and dry panning techniques for use with gold ores in arid regions such as Western Australia. The present day air table was developed in the last century for this purpose. Devices and Processes Classification. All classifiers make a separation on the basis of size, shape, and specific gravity. In addition to specific gravity and size, particle shape factor is often particularly important to pneumatic processes, and advantage is taken of the fact that both flat and fibrous particles settle at velocities substantially lower than that of their equiv¬alent spheres. Another factor of importance in this process is the bulk density of the material since substances such as exfoliated ver¬miculite or partially opened fibrils of asbestos have bulk densities significantly lower than those of the pure in situ mineral. Winnowing takes advantage of all of these physical factors to affect a separation starting with a closely sized feed, and vermiculite has been separated from rock by this process.142 When air classifiers are used to size minerals, the separation will often not be perfect because heavy particles will respond similarly to somewhat coarser light particles while flat, elongated, and low bulk density particles will act as would particles finer than the cut size. An example of a classification process that also results in concentration is the zig-zag classifier"' used to separate paper from glass, metal bottles, and cans during the recycle of municipal solid waste. Application of this process to minerals showing the proper specific gravity, size, and shape differ¬ences is obvious. An important device of this type is an air-aspirated screen used to remove asbestos from its associated rock during the dry processing of that mineral.144 The raw ore is crushed and screened and the asbes¬tos removed by aspiration from the surface of the screen by virtue of its shape and low bulk density. Hindered Settling. Based on the discussion in the section on "Hindered Settling Concentration and Jigging," the use of air devices could be anticipated in desert areas or where moisture may be deleteri¬ous to the product to be separated. Such processes have three major defects: (1) dust may be difficult to contain, (2) fines are difficult to process, and (3) the process is inherently less efficient than are wet processes as may be seen from Eqs. 3 and 4 in the section on "Hindered Settling Concentration and Jigging p. 4-47." Taggart142 has reported on the use of both dry panning and a dry rocker for gold separation in and regions. The dry panning makes use of winnowing, defection on the basis of specific gravity and shape during fall, and hindered settling. Prescreened placer material is poured from one pan to another as air blows across the falling stream. The process is repeated many times with hand picking to produce a rough concentrate. Magnetic minerals are then removed, and the residue, one grain deep in the pan, is air blown. In dry rocking the gravel is sized on a steeply inclined screen with the undersize fed to a riffle box with a porous bottom which is blown from underneath."' This device was the forerunner of the present-day air table. Air Tables and Jigs.145. 146 At least nine types of these devices have been developed, all but two of them for coal. While the devices used for coal may be characterized as pneumatic jigs, pneumatic tables, and pneumatic landers,145 only pneumatic jigs survive today. The same is true for ores.'" In these devices presized particles are fed to the separator which consists of an inclined vibrating conveyor with a porous surface through which air is carefully introduced to form a fluid bed. The lighter particles are lifted by the air out of uphill conveyance and float downhill, while the heavier particles in contact with the surface are conveyed uphill by the vibrator. Typically both transverse and longitudinal slopes are used. A plan view of a Triple S air table and the type of separation made is shown in Fig. 44.147 Because of its external appearance, the device has been called an air table, although it functions essentially as a jig. However, some of the asymetrical acceleration conveying function of a standard shak¬ing table is also performed by the eccentric vibrator incorporated into this device. Low pressure air is admitted below the vibrating table surface consisting of cloth (e.g., canvas), porous plastic, woven wire, or punched metal supported on a wire mesh or grid. As with a shaking table, the device produces a concentrate, middlings, and tailings. The middlings are recycled or are subsequently treated on another concentrating device. As with many wet separations, presizing by screening is necessary. The oft-repeated admonition of Professor Richards, "Separation without classification is damnation," is cer¬tainly true here. In this way a small heavy particle which might weigh the same as a large light particle and thus report to the same place on the air table has already been removed in the sizing step. Air tables are almost universally encased with a canopy to remove dust by aspiration. Some concentration may also be achieved in this step when favorable specific gravity, size, and shape differences are present. Applications Ore Separations. Air tables were originally developed for ore separations, and they find application in and regions or where water is deleterious or inconvenient to use or remove (e.g., small tonnage materials already dried). They have been used for asbestos, bauxite, calcite, cassiterite, columbite-tentalite, diatomaceous earth, fluorite, gilsonite, graphite, kyanite, manganese minerals, mica, managite, perlite, pyrite, pyrrhotite, vermiculite, and uraninite ores. More recently, air tables have found favor in separating a wide variety of secondary materials such as abrasive grains, bone char, catalysts, fiberglass, scrap glass, scrap wire from its insulation, prills from slag, dross and coke from metal or from each other, metal from crushed crucibles, lead from plastic in old batteries, and cubic particles from flat ones. The devices also finds considerable application in the preparation of food¬stuffs. As stated before close sizing before separation is desirable, and if the specific gravity difference is slight, very tight

Jan 1, 1985

By Charles W. Rountree

Industrial Engineering in the industrial world is a relatively new field, having been conceived about the turn' of the century by Frederick W. Taylor, who was not satisfied with the hit-or-miss methods of evaluation and estimation. Through crude Industrial Engineering techniques never before used, he completely changed the bricklaying industry so that work accomplished was more than doubled. To help us better understand what we mean by Industrial Engineering, I think a very simple definition Is in order. Industrial Engineering is a scientific approach to problems in cost, production and operations: and is designed to provide a means of increasing producitivity and reducing costs through standards measurement and standardization of the best methods, practices and equipment, There are many complicated definitions, but pure and simply, it all boils down to providing the means for increasing productivity and reducing the cost. Industrial Engineering has been applied in manufacturing since Frederick W. Taylor?s time and in the last 20 years has grown in stature by leaps and bounds. It is a basic function of successful manufacturing companies and is an accepted fool firmly entrenched in manufacturing organizations. You may know of some successful manufacturing companies that do not employ Industrial Engineers as such, but someone in that organization is doing Industrial Engineering work. They have standards and controls whether they call it Industrial Engineering or not. The manufacturing industry has benefited from this tool; other wise, it would have been discontinued long ago.

Jan 1, 1958

By C. J. Dungey

INTRODUCTION The County of Cornwall is situated in the South Western extremity of England, see Fig. 1, and has for hundreds of years been an historical centre of mining activity. In the last century many thousands of Cornish miners emigrated to Canada, U.S.A., Australia and numerous other countries throughout the world to form the nucleii of mining industries in those lands. At present there are four active underground mines within the county, all mining tin, together with a relatively large china clay industry. All the operations are directly or indirectly associated with the three main granite intrusions, namely the Land's End Granite, the Carnmenellis Granite and the Bodmin Moor Granite. The report which follows relates to two mines, owned by the South Crafty Group of Companies, namely the South Crofty Mine itself and its smaller subsidiary, Wheal Pendarves. Both these mines are adjacent to the Carn Brea Granite ridge, see Fig. 2, which is an off-shoot of the main Carnmenellis Granite mass. The Carn Brea Granite has provided one of the most intensive areas of copper and tin mineralisation anywhere in the world. Many of the mines worked during the last two centuries and the early part of this century were situated close to the metamorphic aureole surrounding the granite intrusion. Both South Crofty and Wheal Pendarves have their shafts collared in the overlying slate, locally known as killas, but the major part of their workings are in the granite beneath. The ore bodies, or lodes, of both mines are generally steeply dipping and approximately one to four metres in width. They have been, and continue to be, worked by either back or under-hand stoping with the major level interval being approximately 37 metres. A feature of both stoping systems is that they require no back filling and in many cases the completed working areas have remained open for many years due to the competent wall rock conditions. The stoping operation necessitates having box-holes or draw points at regular intervals along the strike length at 4 to 7 metres centres; thus a stope some 50 metres long may typically have ten box-holes or draw points through which air can percolate into the empty stopes above. The granite country rock contains traces of uranium minerals, namely pitchblende or uraninite, and it is the decay of this uranium mineral which leads to the formation of the mobile radon gas which, together with its short lived daughters, may create a hazard. The presence of uranium minerals has long been known, but the quantities being so minute, it was not appreciated until some ten to fifteen years ago that these might present a hazard. The Mines and Quarries Inspectorate approached Management approximately ten years ago with a view to taking samples in areas of the mine that they felt might contain the highest concentrations of radon. Samples were thus taken in several areas of both mines, normally well away from working areas and where there was no through ventilation. The results of these initial tests were published by the National Radiological Protection Board in a paper entitled "Radon and Its Daughters in Various British Mines". The results of the initial surveys made it apparent that improved ventilation was the most important factor in reducing radon daughter concentrations within the mine atmosphere, with increased flow rates having three beneficial functions, viz. (1) To dilute radon daughter concentrations. (2) To prevent radon and its daughters reaching a state of equilibrium. It is not uncommon for radon and its daughters to have a disequilibrium factor of 1,000 in some uranium mines.

Jan 1, 1981

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

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

Jan 1, 1982

By Richard L. Breithaupt

INTRODUCTION A shuttle car is a self-propelled rubber-tired haul¬age vehicle designed specifically for underground min¬ing, primarily in coal mines (see Figs. 1 and 2). The drive mechanism is located on the sides of the vehicle, and the load is carried in the center. A chain-and-flight conveyor distributes the load when the car is being loaded by a loading machine or continuous miner and also discharges the load onto a conveyor belt or into a mine car. As the name implies, the shuttle car shuttles back and forth between the working face and the un¬loading point and is not required to turn around. Shuttle cars have served as primary haulage in this country since 1940 and still perform this function in most coal mines today. TYPES OF SHUTTLE CARS Shuttle cars may be battery-powered, diesel-powered, or cable-reel electric powered. Battery-powered cars were used between 1940 and 1950, but are almost out of existence today because of their limited power. Diesel cars have been used in coal mines, but are used mostly in noncoal mines where rules and regulations are not as stringent. Also, objections by organized labor have limited the use of diesel-powered shuttle cars in coal mines. Today, the largest number of shuttle cars are cable-reel cars. Electric power is fed through a cable stored on a cable reel that may have a capacity from 152 to 244 m (500 to 800 ft), depending upon the size of the cable and the cable reel (see Table 1). This, of course, limits the travel of a shuttle car. However, time studies have shown that for the most efficient haulage cycle, the length of haulage should not exceed 91 m (300 ft). As haulage distance increases, the waiting time for the loader or miner becomes longer, and haul¬age efficiency suffers. Cable-reel cars are manufactured in standard hand and opposite standard hand models. A standard car has the operator's control on the left front side and the opposite standard car has the opera¬tor's controls on the right front. Generally a loader or miner operates with two shuttle cars, one of each type. This allows the trailing cable to lie on either side of the entry and prevents the cables from fouling each other. Shuttle cars are designed to operate with either dc or ac power. Cars using dc power are designed for 250-v supply and ac cars are designed for either 440, 550, or 950 v, three-phase voltage. Most coal mines that are relatively level can use dc cars where dynamic brak¬ing is not required. For mines having steep grades (in excess of 15%) ac cars are used because ac traction motors have inherent dynamic braking and are safer when tramming downgrade. A recent innovation in shuttle car design in the last five years has been the introduction of an ac/dc car. This design utilizes ac input power through the cable, and the power is rectified on the car to 250-v do for traction motors. The pump motor and conveyor motor are powered by ac voltage. The main advantage of this design is efficiency-a much smaller cable may be used and the smooth operating characteristics of the dc traction motor is obtained. This prevents shock loading on the traction drive train, which is an undesirable feature of ac traction motors. SHUTTLE CAR FEATURES Elevating Discharge Today, all shuttle cars are equipped with an elevating front conveyor that allows the shuttle car to discharge into a mine car or belt feeder without the use of ramps. Hydraulic Power System Today, all shuttle cars are equipped with an hy¬draulic power system. An hydraulic pump is driven by an electric motor which provides hydraulic power for steering, elevating the front conveyor, and powering the

Jan 1, 1982

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

GENERAL STATEMENT The primary objectives of any field data gath¬ering effort should be to (1) identify and gather the data necessary for the project and (2) obtain the data in a state-of-the-art manner. All too often the initial field data are collected both areally and tem¬porally in an illogical manner without the guidance of a conceptual model of the ground water flow systems involved or even a review of existing geo¬logic literature on the area of interest. The initial data collected frequently are of limited value while necessary basic reconnaissance information is miss¬ing. Initial field data should be collected with the intent of developing a hydrologic overview of the potential mine site and surrounding area. Ob¬viously, one of the initial objectives is to define the area requiring a hydrologic investigation. The data requirements should be identified by the time frame in which collection should be made and by the corresponding increase in sophistication of the data requirements with development and operation of the mine. The data requirements are summarized in Table 1. INITIAL LEVEL SITE INVESTIGATION Area Determination The initial task of any hydrogeologic investi¬gation is to determine the boundaries of the area requiring study. Obviously, the site of the proposed mine is included in the study area. The areal extent beyond the site may be determined from an eval¬uation of existing geologic and topographic maps. Those formations that overlie the ore body, the formations containing the ore body, and the formation(s) that lies immediately beneath the ore body are of direct concern for proper site recon¬naissance. Additional formations below the ore body may require study depending upon their thick¬ness, hydraulic conductivity, and degree of inter¬connection with the mine workings. This initial viewpoint identifies hydrostratigraphic units based strictly on geologic concepts such as mineralogy and structure. Formation outcrops, synclines, an¬ticlines, faults, and fracture and joint patterns are used to delineate the area of the site reconnaissance. The simplistic hydrogeologic environment (il¬lustrated in Fig. 3, chapter 2) requires that field data be collected via test wells and/or geophysical techniques. This approach is necessitated by the lack of surface features such as formation outcrops, streams, and springs. Fig. 5 (chapter 2) illustrates a slightly more complex hydrogeologic regime. The potential mine sites at locations A, B, C, D, and E each intercept a different ground water flow sys¬tem or combination of flow systems. Therefore, each mine location requires that a different area and size of area be investigated. A more complex geologic setting as illustrated in Figs. 6 and 7 (chapter 2) may be approached differently. The area included for the site recon¬naissance should encompass sufficient surrounding area to include the outcrops of those formations suspected of being influenced by the future mine. Even adjacent areas not suspected of being influ¬enced may be investigated if the formations of in¬terest crop out in those areas. Such an extension of the area of investigation would provide a greater regional understanding of the hydrogeologic properties of the formations (hydrostratigraphic units) of interest. Geologic Investigation The initial step before conducting the site re¬connaissance is to review all existing literature on the geology of the area. Existing information should be augmented with new exploration data on the dip, strike, thickness, and lateral extent of the for¬mations in the area. Exploration hole logs should be reviewed for indications of lost circulation, rub¬ble zones, and water producing zones. Existing aer¬ial photos such as those available from the US Department of the Interior, EROS Data Center,

Jan 1, 1986

By William L. Jude

INTRODUCTION Although the mining method described in this chapter is locally designated "modified Mitchell," there are so many features which are characteristic of square sets that the designation "modified square set" would describe the method equally well. The method was developed as a result of local operating experience with various mining methods, observing operations in mines in the southwest where Mitchell stoping had been thoroughly developed, and an exhaustive study of min¬ing methods literature. Since 1941, modified Mitchell stopes have con¬tributed over 50% of the total stope production and were largely responsible for supporting production dur¬ing the war years and period of declining manpower and increasing costs. GEOLOGY The community of Gilman is situated on Battle Mountain at an elevation of 2743 m (9000 ft) in the Battle Mountain mining district, about 45 km (28 miles) by highway north of Leadville, CO. The sur¬rounding area, including the Mount of the Holy Cross and Eagle River Canyon, is noted for its beauty and recreational facilities, both summer and winter. Massive sulfide ore bodies of the limestone replace¬ment type occur as chimneys and mantas or as blankets in strata dipping 0.19 rad (11°) northeasterly. A chimney is shaped like an inverted cone with shale, quartzite, and porphyry caprocks forming its base. This relatively impervious capping apparently blocked rising ore constituents which formed the chimneys and mantas. At the apex or lower extremity of the chimneys, the ore follows small fractures and bedding planes in under¬lying quartzite. Dimensions of chimneys vary, ranging up to 61 m (200 ft) vertically and 76 m (250 ft) horizontally near capping. The manta or blanket ore bodies vary in vertical cross section from one hundred to several thousand feet. One such manta was developed from its outcrop to its chimney source approximately 610 m (2000 ft) downdip. Another manta of substantial cross section and length did not follow shale but extended updip with limestone capping from its source at a lower horizon. All chimneys, however, make up to the shale¬quartzite-porphyry capping. Chimney ore minerals are mainly pyrite and chalco¬pyrite with accessory copper and silver minerals, surrounded by, and in places overlain with, marmatite, pyrite, and galena. Manganosiderite extends into the ore. It may or may not be impregnated with sufficient marmatite and galena to be of commercial grade. Ore may be massive, coarsely crystalline, or sandy, some¬times rubbly. Generally speaking, chimney ore masses are incompetent and of variable grade. Caprock is heavy and caves readily. Manta ore minerals are mostly marmatite, pyrite, and galena associated with manganosiderite. Structural characteristics of the ore are similar to those found in chimneys and where the manta contact shale capping the ground is unusually heavy. However, the large manta with limestone capping mentioned previously consisted of a hard strong ore mass in fairly strong host rock. Commercial ore contacts are fairly regular in this manta but in others the ore waste contacts along the walls are irregular with shaly limestone pendants extend¬ing downward into the ore at capping. Dolomite sand is also quite frequently encountered along wall contacts and less frequently overhead. MINING METHODS Practically all stoping in the Gilman ore bodies has been carried on with timber and fill methods and the features which have been briefly described have dictated the choice. Although at the present time modified Mitchell and underhand square-set stopes account for the total production, considerable production has been obtained from alternate horizontal cut-and-fill stopes with Mitchell truss stopes in the pillars. This method was highly satisfactory in the hard manta ore body mentioned earlier and where ore excavations 11 m (35 ft) wide and 23 m (75 ft) long required no auxiliary timber support. It was a pillar recovery application of the method devised by M. W. Mitchell at the Calumet

Jan 1, 1982

By Erling Stranden, Leiv Berteig

INTRODUCTION The measurement of the total activity of radon daughters in the air of mines has become a routine procedure in order to control the radiation exposure in miners due to the inhalation of these radionuclides. Normally the measured concentration is given in terms of total potential L-energy of the short lived radon daughters. The unit most frequently used is 1 WL (working level), which is defined as the potential a-energy of 1.3.105 MeV per 1 of air. The a-dose to the respiratory tract is, however, not proportional to the inhaled potential a-energy. In addition, the degree of equilibrium between the daughter products in air and the fraction of daughter products not attached to aerosol particles (the unattached fraction) must be known. Theoretical calculations of the equilibrium between daughters and the unattached fraction have been presented by different authors (Jacobi, 1972; Porstendörfer et al, 1978). Such calculations must take into account the plateout of daughters on surfaces, the attachment of daughters to aerosols and removal of daughters by the ventilation. The situation in mines will however be very complex and no single model will cover all situations. In most cases the radon and radon daughter concentrations will be described by a combination of different models. Domanski (1979) used the experimental results of Breslin et al. (1969) from uranium mines to relate the equilibrium between the individual daughters to the equilibrium factor, F. Once the equilibrium factor is known, this seems to be a useful approach to assess the individual daughter concentrations. The concentrations of radon and daughter products may vary considerably during the day. Seasonal variations are also frequently found. It is therefore important to have knowledge of the magnitude of these variations and of the factors having the strongest influence upon the concentrations. The unattached fraction of radon daughters in uranium mines was investigated by George and Hinchliffe (1972) and Raghavayya and Jones (1974). Mercer (1975) later modified the results of Raghavayya and Jones. Both these investigations indicated that the aerosol particle concentrations was the main influencing factor upon the unattached fraction. In this paper we will summarize the main results of a study on the radiological characterics of nonuranium mines. We will also discuss the correlations between the unattached fraction of the potential a-energy and the unattached fraction of the individual daughters, and between the equilibrium factor F, and the individual daughter ratios. More detailed discussions of this study will be published elsewhere (Stranden and Berteig, 1981 and Stranden and Berteig 1981 a). MATERIAL AND METHODS During 4 two-weeks periods corresponding to the seasons in the Norwegian mine where the highest radon daughter concentrations occure, the parameters listed in table 1 were studied. TABLE 1 Parameters measured during the study [Parameter Methods Radon Scintillation flask/Continuous ionization chamber Radon daughters Tsivoglou(1953)/Kusnetz(1956) Respirable dust Cyclone Unattached fraction Wire screen (Stranden and Berteig 1981) Atmospheric pressure Continuous baragraph Temperature and Continuous thermohygrograph humidity] At a fixed location, the radon concentration was measured continuously by a ionization chamber (Stranden, 1981), during these two-weeks periods in order to study short-time and seasonal variations. RESULTS Correlation between equilibrium factor and individual daughter ratios. The mean value of the equilibrium factor in all Norwegian mines measured during the last ten years was 0.46 with a minimum of 0.08 and maximum of 0.93 (Stranden and Berteig, 1981). The correlations between the equilibrium factor F, and the individual radon daughter/radon ratios were studied and the results indicated high correlations. In table 2, these correlation formulae are listed together with the correlation coefficients. TABLE 2 Correlation between the equilibrium factor, F, and individual radon daughter ratios [Formula Correlation coefficient RaA/Rn = 1.14 F 0.551 0.76 RaB/Rn = 1.00 F 0.940 0.99 RaC/Rn = 0.77 F 1.120 0.92]

Jan 1, 1981

By J. L. Kohler

The knowledge-based expert systems (KBES) is a potentially powerful technique to the mining engineer in traditional areas of mine design and operations, as well as in the emerging areas of mining automation and robotics. Since KBES is still experimental in many respects, caution must be exercised as this "high tech" tool is developed and applied to mining. The objective of this paper is to demystify expert systems by introducing the elements of KBES, discussing the strengths and weaknesses, and identifying some areas of application in mining, based on the authors' experiences.

Jan 1, 1986

By L. A. Pursell

The case history, which is the subject of this paper, involves the expansion of a lime plant in Texas. The major new production unit is to be a 545.45 tonne per day coal fired rotary kiln. Auxiliary facilities include additional quarry haulage equipment, modifications to the crushing plant, a hydrator, storage facilities, and a loading system. The existing plant consists of a high calcium limestone quarry, two coal fired rotary and 10 gas fired shaft kilns, nth crushing, screening, stockpiling, hydrating, storing and loading facilities. The proposed modification would increase quarry volume from 545,000 tonnes per year to 805,000 tonnes per year. The plan also includes the addition of dolomitic lime to the product line which requires the purchase and handling of 186,000 tonnes of screened, washed dolomite yearly. A decision to undertake such an expansion was previously a simple and straight forward process. Such a decision now faces a major and complicating obstacle which can delay the project, raise the expected cost, or prevent it entirely.

Jan 1, 1980

By C. N. Speltz

The cost of transportation of industrial minerals is frequently higher than the cost of the product. Where a minc, industrial mineral plant. or customer is served by a single railroad, that single railroad, because of lack of competition, may increase rates and/or provide inadequate service so that the economics of that mine, plant or customer are adversely affected, or so that the level of service precludes satisfactory performance. Transloading of industrial minerals from rail-to-truck or truck-to-rail in order to improve economics and sales of product is not new. However, transloading to reduce freight costs or improve inadequate service caused by a lack of competition in the rail industry, is a relatively new concept brought on by the recent mergers of many of the main line railroads, both in the eastern and western United States. For example, with completion of the Burlington Northern/Santa Fe and the Union Pacific/Southern Pacific mergers, the western United States has only two mainline railroads. The industrial mineral producer, in order to reduce the cost of transportation, or increase the level of service, must evaluate potential transloading opportunities in light of higher prices and reduced service brought on by reduced railroad competition.

Jan 1, 1998