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By Lewis L. Oriard

INTRODUCTION In recent years, there has been a trend in the direction of larger drilling equipment and larger diameter blastholes. Although this change has improved the efficiencies and reduced the costs in many operations, it has increased the potential for damage to underground openings. In addition, in many instances one now finds more sophisticated delicate instruments, automated control facilities, and a large variety of structures in proximity to blasting activity. The combined effect of larger-scale blasting activity and its proximity to various features of interest is such that there is an increased need for a more refined analysis of blasting effects and their control. BLASTING EFFECTS ON ROCK SURFACES The Breakage Mechanism In order to develop techniques for controlled blasting, one must first understand the features of the mechanisms by which blasting causes rock breakage to occur. These features have not been easy to demonstrate, mostly due to the difficulty in making tests and observations at the high stress levels and short time durations involved. When an explosive charge is detonated, the material surrounding the charge is subjected to a nearly instantaneous, very high pressure [on the order of 1.4 to 13.8 GPa (0.2 to 2.0 X 106 psi), depending on the explosive]. If the charge is coupled to "average" rock, this pressure will pulverize the surrounding rock for a distance on the order of 1 to 3 charge radii in hard rock, and to a greater distance in softer rock (this is also dependent on the type of explosive). As the pressure wave passes into the rock, high tangential stresses cause radial cracks to appear, and the nearly discontinuous radial stress zones gen¬erated by the shock front may cause tangential cracks to appear. The extent of these cracks depends on the energy available in the explosive, how quickly the energy is transmitted to the rock, and the strength properties of the rock. The discontinuous shock front is quickly dis¬sipated, but the expanding gases generate a longer-acting pressure. A compressive pulse travels to the nearest face or internal rock boundary where it is reflected in tension. The tensile strengths of most rocks are roughly 40 to %o of their compressive strengths, so the rock may now fail in tension whereas it may have been able to support the diminished compressive phase without failure. The ten¬sile deflection typically produces a failure described as tensile slabbing or scabbing. Laboratory experiments and field experience have pretty well established that several mechanisms are involved. These include (1) the classical case of tensile parallel slabbing when the pressure pulse is reflected at a free surface; (2) failure under quasi-static compressive loading (the shape is normally irregular due to discontinuities in the rock); (3) radial cracking under the action of tangential stresses at the periphery of the expanding pressure pulse; (4) peripheral cracking at the discontinuous shock front which is quickly dissipated; and (5) additional mass shifting due to the venting of the explosive gases. The first three items have received much attention in the laboratory and the literature. The complex effects of gas venting are difficult to test in the laboratory because of the difficulty in reproducing the many weak planes and discontinuities typical of most field conditions, which play such a prominent role in determining the behavior of the rock mass subjected to blasting. Unfortunately, gas venting effects can be pro¬jected to significant distances under certain field conditions, and are sometimes difficult to control. It is not unusual for gas venting to be the overriding factor in determining the final geometric shape and physical condition of the finished excavation. Sources of Damage For the purposes of this discussion, damage includes not only the breaking and rupturing of rock beyond the desired limits of excavation but also an unwanted loosening, dislocation, and disturbance of the rock mass the integrity of which one wishes to preserve (such as mine pillars, underground openings, etc.). The sources of damage include, of course, all those physical features of the rock breakage mechanism. Each of these effects must be limited to the desired zone of breakage and excavation if the integrity of the remaining rock mass is to remain undiminished. The primary zone of rock breakage usually can be controlled in the normal process of field experimentation to determine proper charge sizes and location for primary excavation. However, it frequently happens that there is damage from sources which are more difficult to account for in the design process, which are often overlooked. These are (1) the overbreak due to poor drilling control, (2) dislocation of rock (mass shifting) due to venting of explosive gases, and (3) loosening or dislocation due to the influence of seismic waves (ground vibrations). CONTROL OF ROCK BREAKAGE Importance of Geometry In studying the rock mass and blasting design con¬siderations, it is important to keep in mind the geometric relationships among charge size, shape, and position, and the physical features of the rock mass to be preserved. The features of principal interest are the external shape and position of the rock mass relative to blasting, and the position and attitude of any weak planes in the rock mass. The Sequence of Blasting and Excavation Events Unfortunately, there are too many times when the task of preserving delicate rock is considered hopeless, and because of this attitude, no further effort is ex¬pended towards caution or control. In such cases there is often a failure to recognize the importance of the se¬quence of the procedures. Attention to this can greatly reduce unwanted effects at minimum cost. Perimeter Control The requirements for perimeter control are highly dependent on the special needs of each particular proj¬ect. The desirable degree of control is a highly variable

Jan 1, 1982

By Thomas D. Boyce

Minewater from many western U.S. uranium mines often contains radioisotopes in concentrations higher than permissible for discharge to unrestricted areas. .Radium 226 has the lowest permissible concentrations of all radioisotopes and usually greatly exceeds this limit in untreated effluents. This radionuclide is derived from the natural radioactive decay of uranium: [ ] The treating of minewater effluents has been studied previously (1) and the classical removal method for radium has been adsorption on barytes or freshly precipitated barium sulfate. This paper describes the removal of radium from minewater by utilizing a selective complexer based on a crosslinked organic polymer matrix.

Jan 1, 1982

By Anil Mahyera

Towards the goal of achieving uniform-high-tensions in roofbolts to provide superior roof support[1-7] for extended periods, Terra Tek installed mechanical-anchor bolts in a mine in eight different ways -- under torque/thrust control with and without hardened washers, belleville washers and tension equalization; in a conventional manner with and without a hardened washer. The various bolt groups were analyzed in terms of their tension uniformity and mean tension levels. Installing the bolts accurately with ? a hardened washer under the bolthead, and ? proper torque/thrust control resulted in the mast uniformly tensioned bolts -- with a coefficient of variation of only 14%, an improvement of 3 times over a conventional installation. Also, these bolts maintained a high tension level -- 80% of the installation tension, compared to 50% achieved conventionally.

Jan 1, 1982

By Paul R. Thomas

The Bureau of Mines evaluated the impact of the depletion allowance upon the economic availability of copper from domestic deposits. The potential long-run percentage depletion deduction, cumulated over the life of the properties, was estimated at approximately $2.5 billion for producing properties and up to $11.7 billion for nonproducers. The income transfer, in the form of higher Federal income tax payments if the allowance were repealed, was estimated at $1.5 billion for producing properties and up to $8 billion for nonproducers. The percentage depletion allowance, through its role in determining Federal income tax liability, was found to be very significant in affecting profitability. Repeal of the allowance could result in major domestic copper availability reductions at all price levels, now and in the future, as a result of reduced profit margins for current producers, and postponement of development decisions for nonproducers. It was demonstrated that a 50-percent reduction in the percentage depletion rate applicable to copper would also result in a significant decrease in profitability for nonproducing properties. For producing properties, the percentage depletion rate on copper was shown to be less significant than the income limitation that is placed upon the total amount of the depletion deduction. Repeal of the minimum Federal tax provision was shown to be significant in enhancing profitability for both producing and nonproducing deposits. The report concludes that the percentage depletion allowance operates as an important method of capital recovery and production incentive.

Jan 1, 1982

By Clifford H. Summers

This report describes an experimental program which we conducted on behalf of the U. S. Bureau of Mines, to select from currently available materials two novel systems of construction for air-restraining bulkheads to be used in control of radon-polluted air in underground uranium mines. The program included both laboratory and field tests, the latter conducted in a portion of an operating mine owned and operated by Union Carbide Corporation at LaSal, Utah. The test site was a worked-out, double entry stope having a Volume of about 2000 cubic meters, and was located immediately adjacent to an exhaust airway. The bulkheads were constructed of 15 cm x 1.5 cm timber uprights, 5 cm x 15 cm horizontal members supporting a close-fitted 13 mm plywood curtain-wall., which was covered with a low-permeability membrane to seal in the radon-polluted air. The two membrane materials selected for testing were chosen from a list of twenty-two potentially acceptable candidates using criteria such as strength, flammability, ease of application, adhesion, radon permeability, and elasticity. Thirteen of the candidates were tested in our laboratories for adhesion to plywood and to both wet and dry concrete (simulating wet and dry mine rock surfaces), and for their general physical properties, including radon permeability. The final selections made by the Technical Project Officer, and based on our recommendation, were a pre-formed ethylene propylene diene monomeric rubber (EPDM) membrane 1.12 mm thick, normally used as a roof sealant in industrial construction, and available from Carlisle Tire and Rubber Co., Division of Carlisle Corp., Carlisle, PA, and Aquafas 48-00, a water-based mastic used in a wide variety of commercial applications and available from H. B. Fuller Co., Foster Products Division, Spring House, PA. The EPDM membrane was applied as a single sheet, laminated (dry) between two layers of plywood, and the Aquafas was applied both by troweling and spraying onto a plywood surface. In the field test, radon concentrations were continuously measured before, after, and in the trapped air behind each bulkhead. Temperature, pressure, and velocity of the air in the adjacent airway were also measured and recorded continuously. Radon fluxes at the rock/air interface and through the bulkheads themselves were measured manually on a frequent basis. All continuously acquired raw data were recorded automatically on two USBM Data Aquisition Systems (DAS), and the raw data were converted into engineering units and recorded on magnetic disks by a specially programmed microcomputer that was interfaced with the DAS's. Relatively clean (i.e. low radon concentration) air was pumped into a point near the face of one test bulkhead from a location 170 meters away through vent bag. Several air curtains and two air-restraining bulkheads were erected in passageways adjacent to the test area in an effort to control both velocity and radon concentration of the air flowing past the bulkhead faces. An average concentration of approximately 500 pCi/liter at a flow of 550 m3/minute was attained as the background concentration against which leakage through the bulkheads was measured. The entire field test period extended from October, 1981 to March, 1982, a total of six months, and was subdivided into seven successive time periods: 1) site preparation 2) background measurements prior to bulkhead construction 3) passive restraint, with bulkheads in place and totally sealed 4) open bleeder period of natural convection ventilation in and out of the bulkheaded area through a 10 cm (diameter) opening, with connecting tubing; leading into an Exhaust airway 5) open bleeder period, having forced exhaust through a 1.0 cm opening using a servo-controlled fan 6) charcoal trapping period, using the servo-controlled fan to exhaust air through the charcoal traps, to test the effectiveness of such traps in removal of radon from the exhausted air 7) site decommissioning Each of the two systems tested, one incorporating the pro-formed EPDM-rubber membrane and the other utilizing Aquafas 48-00, was assessed for effectiveness in controlling radon-polluted air by observing the change in the radon concentration in the ventilation air flowing continuously over the surface of each bulkhead. The sensitivity of the observed measurements was limited by the volume flow-rate of the ventilation air and the radon concentration in that air. The data would indicate that the observed increments in radon concentration in the ventilation air moving across bulkheads B-l and B-2 were essentially zero within the limits of sensitivity of the measurements. Data from radon flux measurement taken both on the sealed ribs adjacent to the bulkheads and on the bulkheads themselves did not differ significantly from blank data.

Jan 1, 1982

By E. Divers

This Bureau of Mines report presents results of an underground evaluation to determine the respirable dust and gas control effectiveness of a combined (push-pull) face ventilation system for coal mines. The system utilizes both blowing and exhaust curtain, continuous miners equipped with radio remote control, and flight conveyors. Bureau of Mines tests showed a 97% reduction in respirable dust concentrations from the usual cab position on the continuous miner to the remote control operator's position. This allowed the remote control operator to be well within, and the usual cab position to exceed, Federal respirable dust standards. Tests utilizing a safe tracer gas, sulfur hexafluoride (SF6), to determine the face ventilation effectiveness of the combined blowing and exhaust system showed that the system also has excellent methane dispersion capability. Where wide entries permit, this combined face ventilation system and use of remote control can be a very effective method for dust and gas control.

Jan 1, 1982

By David A. Heitman

The Buick Concentrator is a highly automated operation which utilizes a process control system for continuous monitoring and control. The process control system development, installation, and application to the concentrator grinding and flotation circuits are presented in this paper. The design and implementation of analog field sensing instrumentation, a continuous on-stream X-ray analyzer, grinding and classfication circuit controls, and a recent digital process control system utilizing a centralized computer are discussed. Emphasis is placed on the project planning and installation of the digital control system including a presentation of future plans for continuing application of computer control.

Jan 1, 1982

By Jan Wolski

Today's rapidly shrinking world requires a common language of measurements, a consistent system which could be used by everyone. The S.I. system of units is generally accepted as such and is already in use by a great majority of nations in the world. The change to S.I. units is also slowly taking place in U.S. engineering practice. The transition period is difficult and long because it takes time to get familiar with the new system, to start "feeling" values in new units. Instruments, formulas, computer programs, survey grids, etc. have to be converted or exchanged. The transition period should be used not only for substitution of units, but also for rationalization of presently used calculation methods. This paper examines the major parameters of mine ventilation from the above point of view and discusses a system which is much simpler and more consistent than the American Unit System still in use in the United States. In October 1979, the Society of Mining Engineers of AIME decided to introduce the S.I. unit system in all its publications including the periodical "Mining Engineering", and since February 1980 the system has been used consistently, not only in SME publications but in all conferences and meetings organized by this Society. The authors have no intention of explaining the details of the S.I. system in general. Any reader can find this information in many available publications and conversion tables, and in particular in the ASTM Standard for Metric Practice, E380-79. A comparison is made between various units used in mine ventilation from the past and those presently in use. This will allow an easy translation from one system to another, important when reading publications from abroad. S.I. and customary U.S. formulae are shown side by side for easier visualization of similarities and differences. Recommendations for the necessary steps in order to switch from American to S.I. units conclude the paper.

Jan 1, 1982

By E. L. Smith

The development of larger size flotation machines, which started in the 1950's and 19601s, is still proceeding. The largest mechanical flotation machines available today have unit volumes in the range of 38 to 42 m3 (1,350 to 1,500 cft). In fact, the motivating forces of increasing capital and energy costs and decreasing ore grades, which created the need for large flotation cells, not only still exist, but are becoming more critical. This paper reports the development of an improved mechanism for these large flotation machines. The starting point in the development of a new mechanism was the identification of design and performance features of existing machines where improvements were desirable. The result of this work was a mechanism which is an improvement over existing designs. This new design is characterized by an efficient overhung stator and a vortex profile rotor. The overhung design of the stator removes unnecessary and undesirable baffling in the lowest parts of the tank, while maintaining a quiescent froth layer. This results in lower power consumption and better pulp circulation and solids suspension. The vortex profile rotor acts as an efficient pump, with higher pulp recirculation rates as evidenced by the increased air dispersion capabilities, good solids suspension, and low power consumption.

Jan 1, 1982

By J. Iannicelli

Spray drying is the principal drying method used in the production of over six million tons annually of beneficiated kaolin in Georgia. Thirty-three dryers in the Georgia Kaolin belt comprise the highest concentration of large tonnage spray dryers in the world. The steady growth of spray drying in the U.S. kaolin industry is paced by advances in design and development of large spray dryers which have led to large scale new applications in mining and related industries.

Jan 1, 1982

The disasters between 1920 and 1979 in coal mines in Queensland and New South Wales due to underground fire, gas or coal dust explosion or to gas or coal outbursts, have been assessed. The evidence relating to disasters caused by gas explosion points strongly to divergence from safe mining practice as the main cause. There is a need for greater training and education of personnel, closer analysis of gas reports, better monitoring and testing procedures, improved ventilation standards and a policy where safe working practice is actively encouraged. Conditions in the New South Wales occurrences of gas explosion were similar to those at Collinsville, Queensland in that dykes, igneous sills, burnt out coal and faulting were common to all. Information on outbursts indicates that contributing factors were intrusions, depth of cover, fault zones and other geological disturbances, methods of work, pressure and type of gas.

Jan 1, 1982

By Douglas W. Huber

During the past decade, a general awareness of the need for and subsequent development of increased safety in United States coal mines has taken place. Government and industry alike were stimulated into accelerating their actions through unacceptable accident statistics and costs (in terms of human misery dollar values). Those of us who work in the United States coal industry are aware of the development of governmental controls on nine safety. Both Federal and State regulations have placed increased responsibility upon mine owners and operators to insure a safe working environment and to provide training for miners. No other industrial group of workers has received such attention or, in our opinion, deserves it more.

Jan 1, 1982

By P. M. Daling

This report, sponsored by the U.S. Bureau of Mines, presents the results of an examination of the safety of the impactor shaft sinking system in comparison to conventional shaft sinking technology. Detailed study of the impactor system was also performed to identify and recommend design changes for optimum safety conditions. A comparison of the impactor machine and its support systems with conventional shaft sinking was made on the basis of relative risk levels. Risk incorporates measures of both the frequency and the potential consequence of accidents. A safety evaluation of the component operations of the impactor shaft sinking system was performed using a Failure Modes and Effects Analysis (FMEA) methodology. The results of this analysis show generally that the impactor shaft sinking system is safer than conventional shaft sinking. The potential risk of conventional shaft sinking operations could be reduced by an average of over 35 percent with use of the impactor shaft sinking system. Design change recommendations to further improve the safety of the impactor system include a revised fire suppression system, improved access to and from the operator cabs, reliable noise attenuation measures, and improved communication systems. An operational safety manual was also developed for the impactor shaft sinking system.

Jan 1, 1982

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

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

Jan 1, 1982

By Charles W. Bloomquist

The profitability of miscible flooding in a hypothetical target oilfield is examined. The major costs, including Windfall Prof it Tax, are identified and their re1ative importance are discussed. The sensitivity of profit to important physical and economic variables is evaluated and discussed, including variance in oil recovery, oil price, injectant requirements, injectant cost, and capital requirements. Economic sensitivities to important development and operating decisions such as timing, Spacing, Slug size, and gas recycling are also analyzed. The Windfall Profit Tax (WPT) as it applies to gas miscible EOR projects is reviewed. The effect of WPT on each of the various project cases is analyzed and the effect of reduced WPT on remaining production from existing operations is evaluated. Tax implications are compared for integrated vs. "independent" controlled projects and the alternative of expensing vs. capitalizing tertiary injectants is also evaluated. The impact of the Net Income Limitation in reducing WPT. during early and late project life is determined and discussed.

Jan 1, 1982

By R. Pallini

The Franklin Research Center (FRO, a Division of The Franklin Institute was contracted by the U.S. Department of Energy to investigate the application of elastomeric materials for the purpose of smoothing loads on dragline thrust rollers. This report covers the design, fabrication, and demonstration of a portable- device for measuring roller circle track deviations and roller loadings. The results of this work are documented in this report. During December 1981, a field test group visited the Falkirk Mine, Underwood, North Dakota for the purpose of demonstrating the use of the designed and built device. The data collected indicates the practicality and usefulness of such a device for determining actual dragline roller circle roller loads, while the dragline is in service. The device was useful in determining roller load peaking that may be a contributing Factor in the premature failure of these thrust rollers. The device could be used for field evaluation of elastomeric cushioning materials used to eliminate these load peaks.

Jan 1, 1982

By Milton C. Graves, Marcos Zentilli

"Gold has been reported from most geological environments in Nova Scotia; however, virtually all production (circa 2,000, 000 troy ounces; circa 60 tonnes) has come from that part of the province to the south of the Glooscap Fault which is underlain by Lower Paleozoic rocks of the Meguma Group. Over fifty concordant sets of quartz-carbonate-arsenopyrite veins have seen production. The host rocks are greenschist-grade equivalents of Cambro-Ordovician flyschoid quartz wackes and shales folded and metamorphosed during the Devonian Acadian Orogeny. The veins were emplaced under conditions of greenschist-grade regional metamorphism and under the same conditions of tectonic stress that later resulted in the regional folds and penetrative, axial-plane slaty cleavage. Rapid filling of hydraulic fractures caused by excess metamorphic pore fluid pressure under Acadian tectonic stress conditions can account for the geometry and paragenesis of these veins.The ultimate source for gold deposits in most greenschist-grade domains remains in doubt, but these domains seem to have had ample gold within them available for mobilization. The work done, to date, on the gold geology of Nova Scotia suggests that regional greenschist metamorphism was an effective mobilizer and a potential concentrator of Au, As, W, Fe and S from these vast reservoirs.In rocks of pre-Acadian age north of the Glooscap Fault, gold occurs as an important accessory in volcanogenic massive sulphide ore, as a minor constituent in other types of basemetal concentrations and in one small concordant group of quartz veins.The post-Acadian metallogenetic domain, which overlies and overprints the two older domains, contains small amounts of gold in many mineralizing environments. Veins, with stibnite and gold, cross-cut crystalline rocks. Disseminated gold is present in basal conglomerates of post-tectonic molasse. Gold has been a by-product of Mississippi Valley-type Pb-Zn-Ba-Sr deposits in Lower Carboniferous sedimentary rocks and has been reported with Pb-Zn-Ba-F in veins which cross-cut Lower Carboniferous and older rocks. Small post-glacial placers have also seen minor production."

Jan 1, 1982

By R. O. Burt

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

Jan 1, 1982

By C. C. White

Much benefit can be achieved by a mining company and a manufacturer if both parties can communicate effectively with each other and assist in product development; mutually beneficial ideas and products will develop. This communication can best be achieved if both parties have the opportunity of having on their respective staffs, specialists whose function is to improve the use of products used in mine development. Although a great wealth of roof control knowledge has long existed among mine operators, until recently there was not an abundance of joint development of roof control material by manufacturers and mine operators. This was perhaps due, in past lean years, to the preoccupation of mine management with other overwhelming problems that arise in mining. Manage¬ment had little time for the lengthy process of product development. Until recently, only a few of the largest mining companies had fulltime personnel devoted to roof control. A bolt manufacturer who is invited to work with an operator's roof control people in developing new material is lucky indeed in that he learns what is wanted and he acquires the benefit of the operator's knowledge and experience. He has only to come up with a satisfactory product at an acceptable price - which is not always easy: The development of new products is usually costly and time consuming, and many new products fall by the wayside. In the case of point anchors, however, this product is here to stay. The advantage of point anchors over mechanical anchors is now a proven fact. This advantage is due to less bleed off and higher ultimate load capacity of the resin point anchors. Point anchored bolts have proven to be superior to fully grouted bolts in many areas. We at Birmingham Bolt Company are fortunate in that we have been involved in joint efforts to develop point anchors with several coal companies from whom excellent advice and direction was received. It is appropriate to mention before continuing that the use of point anchors began in the 1960's in Europe and became widespread only recently in the United States. They fall in two categories, point anchors and combination point anchors. These categories have been defined by the Bureau of Mines as follows: (1) point anchors having 24" of grouted length and (2) combination anchors having 24" or more of grouted length. In the present nomenclature, all are anchored with resin or some sort of cementing compound between the rebar or tensioned element and the wall of the hole. The latest development includes an expansion shell as well as rebar, the purpose of the expansion shell being to facilitate installation without the delay necessary in use of types that require a waiting periond for setting of the resin before tightening. (See figures #1 and #2.

Jan 1, 1982

Simple parameters, capable of easy observation on-site under operating conditions without interruption to production, are outlined for the operational assessment of underground conveyor systems. Recording techniques, methods of analysis and deduction of operational characteristics are discussed for batch inputs to several points of a multiple belt conveyor system with limited inter-belt storage capacity. Particular attention is given to input and output load- time distributions, utilisation of inter-belt storage capability and system shutdown behaviour. Actual data, observed for a transfer feeder located between two trunk conveyor belts in an underground coal mine, is used to illustrate the operational assessment of an underground system. Load sensitive feedback systems-are briefly considered.

Jan 1, 1982