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By Terry L. Muldoon

This report describes development and demonstration of an upgraded stench fire warning system for underground wines. Current systems, which have been used since before 1920, release an ethyl mercaptan mixture into the mines' compressed and/or ventilation air during emergencies. Miners, when smelling ethyl mercaptan, evacuate the mine according to predetermined plans. These systems, while simple, have problems with safety and reliability. The upgraded system overcomes these problems. The system uses a different stench agent, is less toxic, less nauseating, and less corrosive than ethyl mercaptan. These combined with the feature of thiophane not reacting with iron oxides, makes the upgraded system safer and more reliable. The system also uses a combination of two simple principles: a standpipe with pressure balancing to provide a constant head, and a sized orifice to control the rate of release of the stench agent. This system enables the mine to release the agent for a longer period which provides better coverage and to ensure that miners will not be exposed to potentially toxic concentrations of the stench agent. The system is easy to operate, requiring only the turning of two 1/4-turn ball valves. The upgraded system has been successfully tested during two evacuation drills at an underground uranium mine. Using three injectors (one on the compressed air line and one on each of two intake air shafts), the system exhibited significant improvements in warning times over the mine's existing system. This report provides a detailed description of the system design and its demonstrated performance in two mine demonstrations.

Jan 1, 1981

By M. Hall, Y. P. Chugh, A. Okunola

Effects of moisture absorption on the behavior of the Dykersburg shale overlying Harrisburg coal seam in Southern Illinois are presented in this paper. Cylindrical samples of the shale 2.5 cm in diameter and 5.0 cm long were studied for weight gain and swelling strains at 90, 95, and 100 percent relative humidity and under immersed water conditions. After attaining equilibrium conditions, the samples were studied for changes in the strength and deformation behavior. The depth of moisture penetration was measured using the Vickers microhardness measurements. Using the basic mass transfer concepts, effective diffusion coefficients were computed for the shale.

Jan 1, 1981

By Tuncel M. Yegulalp

This article summarizes some of the results of the continuing research activity at the Henry Krumb School of Mines, Columbia University, directed towards a better understanding of the stochastic behavior of slurry flows in underground coal-slurry pipeline systems. The slurry system under consideration here replaces all conventional haulage systems between the coal face and the coal preparation plant on the surface. At the beginning of the pipeline system, flow rate and frequency of coal slurry is very closely related to the activities of production equipment at the coal face. Due to several scheduled and unscheduled interruptions in continuous miner's coal mining activity, resulting slurry flows possess stochastic characteristics. Proper understanding of these characteristics is believed to lead to a more realissic and economic design of slurry handling equipment. Fundamental ideas supported by field observations are presented and a Semi-Markov model of the slurry flows is developed. As a result of this formulation, behavior of combined slurry flows from N independent coal faces can be expressed in analytical form. The contents of a slurry surge tank receiving slurry flows with certain stochastic characteristics is investigated as a function of discharge pump characteristics and time.

Jan 1, 1981

By Barry J. Hansen, Ruston A. Ford

Production of mine backfill from the tailings produced in a proposed Crandon concentrator is described. A process has been developed capable of producing suitable fill for the mining method selected for the Crandon orebody. It maximizes the placement of pyrite underground as fill and segregates tailings into pyrite and nonsulfide fractions. To determine the suitability and engineering characteristics of the backfill product for use underground, laboratory and physical stope simulation model testing was conducted. The results of the testing and the application of those results to the design of the backfill system ultimately affecting mine geometry and subsequent mine production will be presented.

Jan 1, 1981

By Tedd B. Bodimer

Why complicate an already complex machine such as a continuous miner by adding roof drills? Immediate permanent roof support is an immediate thought of the mining engineer just as elimination of frequent time-consuming places changes is a thought of the methods engineer. The realization of those two objectives has the potential for improved safety with increased production is the thought of the general manager. Miner-Bolters Used a Generation Ago These are not new thoughts as the 1CM continous miner equipped with two integral roof drills was introduced twenty-five years ago. Approximately two hundred of these machines were sold in the next five years for use in medium and high seam operations. The 1CM miner chassis that remained stationary in the center of an entry or room while the ripper bar was swung and sumped hydraulically provided an ideal platform for mounting two roof drills that could be operated during the continuous miner's production cycle. Two roof bolts could normally be installed in less time than required for the mining cycle. A typical roof drill unit is shown in Figure I.

Jan 1, 1981

By Henry R. Colen

When we were asked to prepare a paper on gold to be part of the Economics of Precious Metals session, our immediate thought was that we should address the question of supply. Are the existing mines of the world capable of filling the demand for gold? We recognized that gold does not conform to the normal pressures of supply and demand. Major new sources of production and increased sales might have a very short term effect and depress the price but this does not last very long. Those forces that affect gold prices are very different from the ones economists normally deal with. Factors such as mass psychology, political tensions and fear are certainly more important to the gold price than supply and demand in today's world.

Jan 1, 1981

By W. Jacobi

INTRODUCTION The radon problem concerns the radiation exposure from inhaled short-lived radon daughters and the associated risk for the induction or promotion of lung cancer. In several groups of miners which were exposed in the past to relative high radon levels during their underground work, an excess lung cancer frequency has been observed. On the basis of these findings it has to be stated that the inhalation of radon daughters is the most important radiation hazard in mining. Involved in this radon problem in mines are the difficulties to measure the individual radon daughter exposure of miners and to estimate the dose to critical cells and target tissues in the lung from inhaled radon daughters. And finally the reduction of radon levels in mines confronts us with special technical problems which are difficult to solve. Thus the setting of annual limits for the radon daughter exposure of miners alone is not sufficient to guarantee an appropriate standard of radiation safety. Of same importance are the improvement of air monitoring and the optimization of technical protection measures in mines. This has to be kept in mind for the following considerations on exposure limits. The ICRP has recommended in 1959 for the first time a maximum permissible concentration (MPC) for the occupational exposure to radon and its short-lived daughters (ICRP, 1959). This MPC in air was derived from a maximum permissible dose equivalent of 0.15 Sv (15 rem) per year to the epithelium of the large bronchi which was regarded as the "critical tissue". This old MPC-value was derived using a simplified lung model which paid main attention to the dose contribution from the inhalation of unattached 218Po(RaA) atoms. Since this time more realistic dosimetric models for radon daughters have been developed; some of these models will be presented in special papers on this conference. Furthermore in the last 10 - 20 years substantial epidemiological information on the lung cancer risk among several groups of Rn-exposed miners has become available. Taking into account these epidemiological and new dosimetric findings the ICRP has adopted in March this year new recommendations on occupational limits for inhalation of 222Rn, 22ORn and theirshort-lived daughters. This report will be published this year as ICRP Publication 32 (ICRP 1981). The recommended limits for radon daughters were derived from the basic dose and risk limits as they were proposed by ICRP in its new basic recommendations (ICRP Publ. 26, 1977). In the following this basic system of dose limitation is shortly outlined before the special recommendations for radon daughters are described. [BASIC CONCEPTS AND LIMITS] The new basic recommendation of ICRP (Publ. 26, 1977) proceed from the assumption that in the dose range which is relevant for radiation protection, the risk RT of so-called stochastic radiation effects (carcinogenic and genetic effects) in a target tissue T increases proportionally with the mean dose equivalent HT in this tissue: RT = aT HT. If several target tissues are irradiated the total individual (stochastic) risk is given by the sum of risks in these tissues: [R = ?TRT = ? T aT HT]. Thus this concept of individual health detriment takes into account the risk contribution from each relevant tissue. It replaces the old concept of the critical tissue, which neglected the exposure of other tissues. Applying this new concept a dose quantity can be defined which is proportional to the total individual radiation risk from stochastic effects. This so-called "effective dose equivalent (HE)" is given by the sum of the weighted dose equivalents in the relevant tissues: [HE = E wT HT with E wT = 1 T '1] The weighting factor wT of each tissues follows to wT = aT/awB, where awB is the total stochastic risk coefficient for a uniform irradiation of the whole body. For the lung as a single organ comprising several target tissues, the ICRP recommends a total weighting factor w = 0.12 (ICRP 1977/79). This concept of risk or detriment, respectively, has important consequences. To meet the objectives of radiation protection the new ICRP-system of dose limitation is based on the following three principles: 1) The justification of practices involving radiation exposures; no pratice shall be adopted unless it produces a net benefit. 2) The ALARA-principle that all exposures shall be kept as low as reasonably achievable; this requires the optimization of the level of protection. 3) The limitation of individual exposure (annual dose limits). It is important to note that the dose limits for individuals are only one part of this system of dose limitation. Furthermore, these limits cannot be regarded as a boundary between safe and unsafe. They have to be interpreted as the lower boundary of an unacceptable exposure region. This means that in addition to the observance of these basic dose limits it is demanded to optimize the level of protection. Consequently the optimized exposure levels or the corresponding operational limits will be normally lower than

Jan 1, 1981

By G Alan Foster

Many complaints generated from blasting operations result from the marked human response to ground transmitted vibrations and air blast effects and so many authorities have considered the regulation of blasting in terms of human response rather than, as previously, in terms of preventing damage to structures.

Jan 1, 1981

By Bruce Pinkston

This report presents a description of the Prototype Demonstration Vehicle developed under U.S. Bureau of Mines Contract No. JO 366003, "Improving Man Transit Safety". This vehicle was constructed to assess the feasibility of a set of equipment related guidelines which were drafted in Phase I of the same contract. State-of-the-Art improvements were embodied in such areas as collision survivability, occupant comfort, and multiple mode (on and off rail) operation. The electric battery-powered vehicle carries a twelve man section crew at speeds up to 12 mph. The description of the demonstration vehicle is presented in Volume I. Safety Guidelines are presented in Volume II.

Jan 1, 1981

By Noel N. Moebs

Studies by the Bureau of Mines have identified geologic structures in mine roof rock that contribute to many roof falls in Appalachian coal mines. These structures, including paleochannels, scours, pinchouts, slickensides, clay veins, crevasse splays, and joints, can often be identified during, and sometimes before, mine development. Mine projections can be revised to reduce the adverse effects of discontinuitites in roof structure; large roof areas of laminated sandstone or incompetent strata generally can be delineated or inferred from exploratory drill hole data, and the need for supplementary support can be anticipated. Accurate descriptions of roof geology also provide some indication of optimum length and type of roof bolts that should be installed.

Jan 1, 1981

Provide respiratory protection for escape in a toxic or oxygen-deficient atmosphere with a device small and light enough to be belt-worn. Approach Combine a 10-minute, belt-worn respirator with a stored, larger, long-duration (60-minute) attachment that can be plugged into the 10-minute unit without removing the mouthpiece. How It Works The 10-minute device will be carried by the miner at all times, and the 60-minute canister or attachment will be stored near the miner. In the event of a disaster, the 10-minute device would be donned immediately. The miner would then proceed to the stored 60-minute unit and plug it into the 10-minute unit. For the 10-minute unit, a breathing circuit was developed in which air is passed through the KO2 bed twice per respiration. The 60-minute canister differs from the 10-minute unit in that air is drawn only once through the chemical bed per respira-

Jan 1, 1981

By Singh BN, Chakravarty PK, Chatterjee LM

Indian blast furnaces operate with coke gener- ally containing 24 to 28% ash compared to 10-15% ash in coke in most other parts of the world. The operating problems are further aggravated by the high alumina content and the adverse alumina to silica ratio in the ore. The above factors not only increase the coke rate but, at the same time, increase the flush slag volume, make the bosh slag highly limey and thereby viscous, and produce a flush slag containing 24 to 26% alumina. Relative- ly low flame temperatures (RAFT) make conditions more difficult. The high basicity of the bosh slag limits the maximum flush slag basicity to around 1.0, beyond which the burden movement is adversely affected by the viscosity of the bosh slag: This effect is partly controlled by operating blast furnaces with high MgO in the flush slag (around.10%), which also improves the desulphurizing power of slag. Steps are underway in some plants to inject lime through theòtuyeres to control the bosh slag basicity and also to neutralise the large quantity of silica which is released at the tuyere level from the coke ash, so that the silicon content of the hot metal can also be controlled. High bosh slag basicity coupled with the characteristics of iron ore/aggolmerates, widens the softening-melting zone and alters the shape of the softening-melting zone. This leads to operat- ing difficulties and poor hot metal quality. This paper discusses the above aspects of Indian blast furnace operation and steps being taken to surmount some of the above difficulties.

Jan 1, 1981

By James R. Jones

The concern for air pollution goes back centuries as will be seen from this quotation : "Strife and coal, it seems, have a hand-in-hand historical relationship. It was thought by some . . . in the Middle Ages that burning coal filled the air with dangerous poisons. So great was the concern that in 1306 King Edward I of England passed the first and unquestionably the harshest air pollution legislation edict that history will record. His proclamation declared the use of coal an act punishable by death." As related to coal mining, the air pollution problem is primarily that of its customer when using the fuel. Thus the major impact relates to markets for coal rather than the actual mining operation itself. In the vicinity of the coal mine the problem is principally related to fugitive dust from roads, transfer facilities, loading facilities, crushers, etc., and particulate matter (dust or fly ash) from thermal dryers and air cleaning tables.

Jan 1, 1981

By C. R. Barker, D. A. Summers, H. D. Keith

Introduction Historically, the presence of methane has been a problem, mainly in and around the working areas of active coal mines, and only in these areas has drainage been considered. Drainage, where practical, has been achieved through the drilling of holes forward into the coal and the surrounding strata from the working area. These holes generaly have been short in length, although where methane drainage operations around a longwall face have been undertaken, the holes have had to be longer in order to adequately drain from the center of the face into the access gate roads. In recent years, attempts have been made to degasify the coal seams in advance of mining, without disruption of the mining cycle. This is done by drilling much longer horizontal holes through the coal in advance of the working area. Under the aegis of the federal government, methods have also been developed for draining coal seams of their methane content in advance of mining, but from shafts sunk from the surface, without using the active area of the mine as the location for the drill holes. Development of methane drainage has recently been encouraged by the potential use of the drained methane as a commercial energy source, with a need, therefore, to adequately organize a collection system, separate from mining the seam for coal. This has already been successfully accomplished, for example, in the Federal No. 2 mine of Eastern Associated Coal Corp. starting in 1975 (Johns). However, whether the system gains access to the coal through horizontal drilling from a pre- existing mine or via access through a separate shaft from the surface, long horizontal holes are required to adequately tap the methane reserve. It is to this regard-the actual drilling of the horizontal holes-that this paper is directed. It will examine potential benefits that may accrue, both in conventional horizontal hole drilling from a mine site underground, and also in drilling from the surface if a high pressure water jet drill is used to drill the degasification holes. Long Hole Drilling from an Underground Site Personnel from the Bureau of Mines have recently examined methods for conventional drilling of long horizontal holes to gain access for methane drainage. They have shown that it is possible (Cervik, Fields, and Aul) to drill out some 610 m using a conventional drilling system. Three types of bit were used in the program and by alternating between a drag bit, tricone bit, and plug bit, advance rates of between 0.6-3.6 m/min were achieved. Hole diameters varied from 7.6-9.2 cm in surface tests at bit thrusts of 1360 kg. A hole was then drilled and maintained in relative alignment within the coal seam for a distance of 640 m. Thrust levels had to be lowered to between 363-680 kg across the bit. Because the loads were smaller than those used in the surface trial, advance rates in the hole were of the order of 10-38 cm/min. The thrust level was lowered since it was found that the level of the thrust controlled the inclination of the drill so that, for example, a thrust of 363 kg caused the hole to incline downward, while at greater than 544 kg the hole inclined upward with the 9-cm-diam bit. Thrust levels increased 227 kg when the hole diameter was raised to 9.2 cm, although in such a case penetration rates in excess of 56 cm/min could be achieved. Horizontal Water Jet Drilling of Coal The University of Missouri-Rolla has recently undertaken research for Sandia Laboratories on the use of high pressure water jets as a means of drilling through coal. The initial experiment in this program called for drilling a hole horizontally into a coal seam from the side of a strip pit using water jets as the cutting mechanism. A very simple setup [(Fig. 1)] was used in this program and a 15-m hole was drilled at an approximate drilling speed of 1.2 m/min. The nozzle was designed so that the hole dimension was approximately 15 cm across [(Fig 2)] and the thrust was maintained at levels below 91 kg in moving the drill into the coal face. The system used was very crude and comprised a high pressure water jet drill enclosed within a 5.7-cm outer diameter galvanized water pipe to provide rigidity to the drilling system. This pipe sufficed to maintain hole alignment over the 15-m increment. While it is premature to make long-term predictions on ultimate applicability of this sytem to long hole drilling, certain inherent advantages of water jets can be delineated from research results and suggest considerable advantage to further research in development of this application. High pressure water was supplied at approximately 62 046 kPa from a 112-kW high pressure pump, with a 83 L/m flow through the supply line to the nozzle. The drilling system consisted of a nozzle rigidly attached to the front end of the galvanized piping. High pressure fluid was supplied to this nozzle through a flexible high pressure hose that fed from the nozzle back through the galvanized pipe to a rotary coupling attached

Jan 1, 1981

By Dan S. Nilsson

When an open pit mine is deepened, more and more waste rock must normally be removed. Techniques for optimization of open pit design are well-established within the mining industry. The recommended method is to choose the open pit alternative with the highest net present capital value of open pit and future underground mining. When the open pit is mined out, the operation must be closed, or, if it is profitable, mining must start underground. When planning an underground operation we experience the same type of optimization problem that occurs when determining the optimal depth for open pit mining. When mining underground the ore body is often divided into a number of areas, blocks or slices, each containing a transportation system with trains, trucks, conveyors, etc. When all the ore within such a part of the deposit is extracted, the operation must be closed, or, if it is profitable, a new transportation must be developed. The question that now arises is: how large a pan of a deposit shall be included in the first development and how much shall be left for future consideration?

Jan 1, 1981

As part of the overall development of a borehole mining system -develop backfilling techniques to reduce possible damage to the environment caused by either the cavities or the piles of sand tailings that result from in-borehole hydraulic mining. (Borehole mining is described in technology News No. 56.) Approach Water filled cavities at the bottom of boreholes are rapidly and economically backfilled by jetting a sand-tailings slurry underwater. How It Works The Bureau of Mines borehole mining system creates a cavity having a horizontal diameter of more than 50 feet. (See photograph). This cavity can be backfilled with much of the sand previously mined. The sand tailings or similar backfill material is dumped by a front end loader into a slurry mixing tank at the rate of about one ton every three minutes. Enough water is added to give a thick, but pumpable slurry. The slurry is pumped down the borehole through a string of 4-inch pipe to the cavity. After passing through a right elbow at the end of the pipe, the slurry is ejected through a nozzle. The resulting slurry jet emplaces the backfill at the outer wall of the cavity. To obtain an even distribution of fill in the cavity, the nozzle is slowly rotated. The nozzle and pipe assembly is supported from a turntable on the surface. See diagram.

Jan 1, 1981

By T. O. Hagan

A system for the monitoring of micrcoseismic events. which is designed for use in feasibility studies of rock-burst prediction, has been operating at Western Deep levels Limited since 1977. Analyses have revealed that the presence and Orientation of geological discontinuities ahead of the mine face can be predicted from the data produced by this system.

Jan 1, 1981

By J. H. Fourie

This paper highlights modern practices in ore-treatment plants, with particular reference to the flotation of P2O5 at the plants of the Phosphate Development Corporation (Foskor) at Phalaborwa in the northern Transvaal. The following aspects are covered: design criteria, equipment and personnel, production standards and control. cost control, maintenance, and process development.

Jan 1, 1981

By Gregg E. Korbin, Robin B. Dill, Corey T. Dare, Tor L. Brekke, Larry R. Myer

A study was conducted to develop a fundamental understanding of the relationship between the size of an advancing tunnel face, the rate of excavation, and stand-up time in squeezing ground. Stand-up time is defined as the time elapsed after excavation that an unsupported tunnel face will remain stable. Squeezing ground is argillaceous rock or soil that exhibits pronounced time dependence of deformation and strength properties in-situ. A case history study documented the nature, causes, and solutions of several cases of stand-up time problems observed in the field. Then, in the laboratory, a series of physical model tests were performed in which model tunnel size, face advance rate, material properties, and pressure (representing overburden pressure) were varied. Stand-up time was found to increase as advance rate was increased or tunnel size decreased. In addition, an influence of advance rate on the definition of stability ratios was established.

Jan 1, 1981

By L. A. Panek

Using an approach based on the theory of similitude, the author develops a general equation and related concepts that provide new insights to an old problem. The load-bearing capacity (strength) of a square, rectangular, or cylindrical prism of a brittle material such as rock, coal, or concrete is expressed as the mathematical product of a size effect, a shape effect, and a function of the mechanical properties (deformability, cleat/joint spacing, and friction) of the pillar, roof, and floor materials. Through an application of multivariate statistical analysis, the results of assorted tests (laboratory and in situ) and observations of failed pillars can be combined into a fundamental scaling equation, which provides a unified, consistent approach to the estimation of mine pillar strength. The form of the equation is sufficiently general to be able to incorporate refinements in the characterization of the structural parameters, as they become available through future analysis and experimentation. Analysis of the published test data suggests that pillar compressive strength (per unit of area) is proportional to the cube root of the pillar width, and that the height effect can be described as the superimposing of a variable size effect (characteristic of the material) on the constant width effect.

Jan 1, 1981