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By Bruce Hebblewhite

"Historically there have been a number of different hypotheses and empirical models developed in an attempt to describe the nature of fracturing above longwall panels in underground coal mining. The motivation for such models varies, ranging from understanding the impact of mining on surface subsidence to back-analysis of caving behaviour in the immediate roof behind the longwall face. One of the most critical motivating factors that is taking on increased importance in many coalfields is the need for better understanding, and hence prediction of the impact of mining on overlying strata, particularly strata units acting as aquifers for different groundwater horizons.This paper reviews some of the major prediction models in the context of observed behaviour of strata displacement and fracturing above longwall panels in the Southern Coalfields of New South Wales, south of Sydney. The paper discusses the parameter often referred to as “height of fracturing” in terms of the critical parameters that influence it, and the relevance and appropriateness of this terminology in the context of overlying sub-surface subsidence and groundwater impact. The paper proposes an alternative terminology for this parameter that better reflects what it is and how it is used. The paper also addresses the potential role of major bedding shear planes mobilised by mining and their potential influence on overlying subsidence and groundwater interference.INTRODUCTIONKnowledge of the detailed nature of rock deformation and failure above any form of large-scale underground mining is always going to be limited to interpretation from a very incomplete set of data. It is extremely difficult, if not impossible, to directly measure the detailed nature of the rock failure, fracture networks, and deformational behaviour above an extracted mining area. Limited techniques such as borehole extensometry can provide some evidence of relative or incremental deformation in the direction of the borehole (usually vertical). However, such data cannot assist below the horizon where full caving has caused major rotation and dislocation of rock blocks and effectively destroyed the instrumentation borehole. Above such a horizon, the data is only valid along the axis and in the direction of the borehole, and to the level of detail defined by the extensometer anchor spacing intervals. It is often also restricted by major shear planes intersecting the extensometer borehole and cutting off the instrumentation below the level of the shear horizons."

Jan 1, 2019

By M. J. Zdunczyk

Sand is one of the most common materials on the earth’s surface. It is a naturally occurring material resulting from the physical and chemical breakdown of rocks. The definition of sand varies, as do the number of products made from the material. Sand is generally classified by grain size, however the measurement or scales differ depending on the context in which it is being used. Geologist define sand as particles of rock between .05 and 2 millimeters in diameter. The variability of the mineral composition generally dictates which sand can be used for different industrial purposes. For example, sand used for Portland cement concrete and asphaltic concrete mixtures can be very different in physical and chemical composition than sand used in the making of fiberglass. In this paper, construction uses of silica sand, which consists almost entirely of the mineral quartz (SiO2), will be discussed and analyzed.

Jan 1, 2001

By Brad C. Lewis

The coal mining industry has experienced a general trend toward longwall operations. An increasing percentage of underground coal is being mined by longwalls. And productivity in longwalls is likewise increasing. Operators of longwalls, however, have not yet reached a state of mining euphoria. Competitiveness within the industry and from alternate fuel sources is forcing continued pursuit of higher productivities and lower costs. [Figures 1, 2 and 3] present trends in the industry. Longwall mines, as defined here, are those reportedly having longwall installations in each of the respective years. Production from both the continuous miner and longwall units is included in the longwall mines. Nonlongwall mines are all other underground mines. Productivity calculations are based on total employee-hours associated with the operation. Annual percent increases in productivity appear to have peaked in the 1985-1987 time frame. This was due, in part, to implementation of technological advancements and improvements in management-labor relations. The results shown in [Fig. 3] aroused the author's curiosity. The first question was "Have technological advancements to increase productivity peaked?" It does not appear so.

Jan 1, 1990

By Stuart R. Cotterill, Jack D. Nelmark

Introduction Opencast surface mining requires blasthole drilling to achieve mine planning. No single method of mining is right for all conditions. Occasionally, mine planners introduce a project with little regard to drilling and blasting. These costs are usually a significant percentage of total mine overhead. One drilling and blasting method that can be used to reduce overall mine expense is inclined or angle hole drilling. But inclined-hole drilling will not cut costs unless a set of required circumstances is met. Several conditions of mine design indicate the use of angle drilling. Geotechnical problems must first be analyzed. Primary considerations are safety, fragmentation, production, and maintenance. Secondary considerations are, in some instances, reduction in amount of explosives, which lowers blasting costs; reduction of practical fragmentation beyond blasting pattern limits; and reductions in overall costs due to reduced volume of excavation. Future trends in blasthole drilling may include incorporation of drill equipment designed to allow rapid change from vertical to inclined drilling. This will allow the operator to adapt to changing hole requirements with minimal production loss. Another improvement may be the ability to chamber holes at any elevation throughout the hole depth. This will allow more charge to be placed at strategic locations to improve fragmentation. Merits of Drilling Inclined Holes Drilling is performed to create voids for explosives. It is done either to recover minerals or to excavate a volume of earth. The manner of mining or excavation can greatly affect the economics of the overall project. Many innovations and methods have been tried, but the most practical and economical is still drilling holes and using explosives to fragment the material. To break up the material, the practice has been to place charges within the mass. This is done by drilling holes of a specific diameter and depth into the material. The holes are of such magnitude as to develop sufficient energy to break or loosen this material, providing for easier removal. In general, more open or exposed faces must exist to aid material breakage when blasting. Inclined drilling can be advantageous in open-cut mining with highwall depths of at least 24 m (80 ft). A minimum of two faces are open in open-cut mining and the material can be thrown outward to cover pre-mined volume. This cuts cost by eliminating the need to remove or relocate a large portion of material. Sometimes the material is of such a mass that more energy per unit is required to loosen it than can easily be done by blasting vertically. Inclined holes allow a more uniform loading of drilled holes. The unit of energy is raised, so little or no secondary blasting is required. This usually comes about only if the main plan calls for drilling holes deeper than 24 m (80 ft). For shallow holes, spacing and burden can be made closer to allow sufficient energy to blast the toe. If the open face is in only one

Jan 9, 1984

By Ivor Hawkes

INTRODUCTION High energy breaking is an alternative to using ex¬plosives in underground secondary breaking operations. It also is a means of upgrading conventional hand-held breakers, manual sledge-hammer breaking, and scaling bar operations. Major areas of application are in sec¬ondary breaking over grizzlies and at drawpoints. Other applications include breaking down ripping lips in longwall seam mining, scaling in stopes and rooms, general demolition work, and roadway maintenance. There is considerable interest in high-energy impact breakers for use in primary ore breaking, but, as of 1977, all such applications have been only experimental (duToit, 1973; Joughin, 1976; Wayment and Grantmyre, 1976). EQUIPMENT Essentially, a high-energy impact hammer is a boom¬mounted pneumatically or hydraulically actuated breaker. The machine basically consists of a piston that oscillates in a housing and impacts the end of a tool or moil thrust against the rock. The force applied to the rock primarily depends upon the impact energy of the piston-the higher the impact or blow energy, the greater the force and, thus, the greater the rock break¬age. Among drill and breaker designers, a common expression for blow energy is "force of blow." Hand-held breakers are limited to blow energies of about 140 J (100 ft-lb), because the operator is unable to handle heavier machines efficiently or to absorb the recoil energy resulting from higher blow energies. How¬ever, these restrictions do not apply to boom mounted breakers; machines with blow energies on the order of 4000 J (3000 ft-lb) and higher are available commer¬cially for underground use. There is considerable evi¬dence to show that increasing the blow energy also in¬creases the efficiency of the breaking operation; i.e., more rock is broken per unit of energy expended (Grantmyre and Hawkes, 1975). Thus, there is a trend to higher blow-energy machines, particularly where high¬strength rocks are to be broken. In relation to rock breaking, the blow rate of boom¬mounted impact breakers is not as important as it is for rock drills. This is because the breaker must be moved over the work surface between blows. The blow rate is governed eventually by the power supply, and typical blow rates range between 200 and 600 blows per minute. As a general rule, light blow-energy machines have higher blow rates than heavier machines. Table 1 lists most of the boom-mounted impact breakers that were available commercially during 1977, and it gives details of the blow energies and machine weights. Restrictions are placed on the blow energy by the machine weight and size, and by the strength of the boom. Typically, boom-mounted impact hammers have a blow-energy to mass ratio of about 1.5, with lower values for lighter machines and higher values for heavier machines. In addition to supporting the hammer weight, the boom also has to absorb the recoil energy of the blow, which can be on the order of 1400 J (1000 ft-lb) for large hammers operating in a horizontal mode. Interesting exceptions to the general run of impactors are the Joy HEFTI hydraulic hammers. In these machines, the piston impacts onto a fluid cushion that is positioned between the piston and the impact tool. This approach allows very high piston velocities, over 30 m/s (100 fps), to be used without the risk of break¬ing the piston or impact tool. Steel on steel impacts must be limited to impact velocities of about 10 m/s (35 fps) due to the high impact stresses generated; thus, increased blow energies can be achieved only by increas¬ing the piston size. The Joy 514 HEFTI®, listed in Table 1, has a blow energy of 27 100 J (20,000 ft-lb), but, as of 1977, the machine has been used underground only on an experimental basis. Using a fluid cushion between the piston and the impact tool allows the use of light pistons, reducing the overall machine weight. The recoil energy, which must be absorbed by the boom for a given blow energy, is directly proportional to the piston to machine mass ratio, and operating with light pistons provides an addi¬tional benefit in reducing the requisite boom size. Both pneumatic and hydraulic hammers are avail¬able commercially. Although hydraulic hammers are a relatively recent development, they already outnumber the pneumatic machines in use. There are many reasons

Jan 1, 1982

By Nicholas T. Camicia

A career in the coal mining industry today offers students one of the most interesting and challenging opportunities available in the industrialized world. As reserves of other fossil fuels dwindle in face of the world's demand for energy, it has become apparent that coal, once the king of energy sources, must be relied upon again as the major fuel for powering industrial growth in developed and developing nations. Centuries of coal supplies lie beneath the earth's surface, assuring the world of a steady, plentiful supply of energy until science and technology can produce a way to harness and efficiently use renewable sources, such as solar. The need to increase our use of coal during the last two decades of the 20th century is reason enough to believe that the future growth of the coal industry is assured. A major impediment to that growth might be backward national public policies that could restrict the full utilization of our coal reserves. These policies could result from overregulation and a failure of the public to recognize that coal use, with proper techno- logical safeguards, is compatible with our society's environmental-quality objectives. Communication of this problem and its solution is a major undertaking of the coal industry in the 1980s. At the end of the 1970s, the American coal industry was producing approximately 635 Mt/a (700 million stpy) of coal. To meet the nation's projected energy needs by 1985, the annual tonnage must be doubled. A similar scenario is predicted for the rest of the world. Such growth in the modern international business world will require huge additional amounts of capital, equipment, and people who know how to wisely raise and apply the capital ; use the equipment productively; and deal with the many community, public, and governmental problems such expansion will create. Whether a student's interest is in actual mining operations, business administration, engineering, finance, law, industrial relations, marketing and sales, or public affairs, a career opportunity is available in the industry.

Jan 1, 1981

By H. Peter Knudsen

In most open pit gold mines, ore-waste selection is based on estimates made from blasthole samples. These estimates frequently are biased. This paper focuses on the source of the conditional bias. While it is true that high grades and "outliers" can induce a bias, they may not be the only reason for the persistent bias seen at many mines. In this paper, it is shown that the poor blasthole sampling can cause a conditional bias in the estimates.

Jan 1, 1993

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

7.1 INTRODUCTION Subsidence measurements include surveys of surface and sub- surface movements and measurements and monitoring of surface structural damages. In this chapter the most commonly employed methods are discussed. 7.2 MEASUREMENT OF SURFACE MOVEMENT Surface movements involve two basic components, subsidence and displacement. The other five components (i.e., slope, curvature, strain, twisting, and shearing) are derived from the two basic components. Since all measurements are geometrical, they can be measured by common surveying techniques. 7.2.1 MONUMENTS The first step is to select the types of monuments or surveying stations. Various types of monuments have been employed. The simplest one is a paint mark on the pavement or concrete slab, although this should be discouraged because it is easily affected by artificial or environmental factors. Others use steel or wooden rods, 2 to 3 in. in diameter or square by 2 to 5 ft long, driven all the way down, flush with the ground. Both types of-monuments have no provision for frozen ground heave and both are tampered with easily. A more elaborate type is shown in Fig. 7.1 (O'Rourke et al., 1982). A rebar or iron pipe, 2 in. in diameter by 6 ft long is driven down to nearly flush with ground surface. The top 4-ft section of the hole around the pipe is enlarged to 1 ft in diameter and grouted with concrete. The most sophisticated type of monument is shown in Fig. 7.2 (Conroy and Gyannaty, 1983a). It consists of two hollow steel pipes. The inner I-in. pipe, 10 ft. long is enclosed by a 2-in. pipe at the upper 5-ft. interval. The gap between pipes is filled with foam to isolate the inner pipe from external environmental effects. The bottom of the inner pipe can either remain open or be sealed with a cap, depending on whether the ground is dry or wet. 7.2.2 MONUMENT ARRANGEMENT AND SURVEY LINES The monuments are installed strategically to form survey lines. Three basic types of monument arrangements are shown in Fig. 7.3. Figs. 7.3a and 7.3b are for single and multiple panels respectively; Fig. 7.3~ is mainly for research purposes with multiple survey lines both in longitudinal and transverse directions. This is extremely useful for defining the distributions of ground deformation, as shown in Figs. 2.10 and 2.11. Basically there are two types of survey lines: the whole basin survey line which extends across the panel width or length (e.g., lines R, R, in Fig. 7.3 a&b), and the half basin survey line which covers approximately half of the panel width or length (e.g., lines R, F in Fig. 7.3 &b). Survey lines are generally straight and linear. They should be located on the major crosssections of the subsidence basin. Survey lines must extend beyond the edge of the opening for a sufficient distance so that the edges of the subsidence basin can be determined accurately. The minimum distance from the edge of the opening must be [ ] where h = mining depth, [ ] = angle of draw, and [ ] = 5 to 7 times monument spacing for whole basin surveys, or 8 to 10 times the monument spacing for half basin surveys. There must be at least two controlling monuments at each end of each survey line. Controlling monuments must be located beyond the influence of underground mining. The spacing of the monuments ranges from 15 to 200 ft. The UK National Coal Board (NCB, 1975) recommended a spacing equivalent to 0.20 of the mining depth. Peng and Geng (1982) recommended that the maximum monument spacing be dependent on mining depths as follows:

Jan 1, 1992

By F. B. Micheli

Introduction and Historical Background Although both sluices and devices using a simple flowing film are among the oldest known methods of concentration, they are still extensively employed in a variety of situations where they are not only efficient but also have a low operating cost. More particularly, they are valuable for making a bulk concentrate of high specific gravity minerals from alluvials and beach sand and in upgrading low value ores. The antiquity of some of the basic concepts can be seen in the fact that they are found in Agricola's De Re Metallica which was published at Basic in 1546, while in 1602, Carew114 gave a fairly clear account of tin ore dressing in Cornwall in which grass turves were used to catch the cassiterite. This is perhaps one of the forerun¬ners of the blanket strake. The planilla of Mexico, the lanchute of Malaysia and the old Cornish box huddle are also descended from these early strakes. In fact, a sluice with holes in the bottom for concentrate discharge was shown by Agricola, not unlike the principle used in the plane table and the lamflo sluice. The convex huddle for the treatment of sandy material appeared about 1848 and the concave form eight to ten years later. For the treatment of "slime" sizes, wooden dead frames which were manually worked were used in Cornwall for many years but gave way to the automatic water-operated type in 1860, and even a mechanically tilted frame was described by Henderson115 in 1857. Theoretical Considerations The behavior of solid particles in suspension depends to a great extent upon the pulp density and the size of the suspended particles. In a fairly dilute suspension, such as that normally used when dealing with small particle sizes, the behavior of particles in a flowing film results from two effects. These are the lateral displacement, which is determined by the time taken for each particle to penetrate the flowing film and reach the solid surface, and the resistance offered by each particle to further displacement after it has reached this surface. The initial penetration through the flowing film depends on the size and specific gravity of the particle and the thickness and viscosity of the film. As a result the smaller particles will migrate further before their movement is retarded relative to larger particles of the same specific gravity. The behavior on reaching the solid surface de¬pends on whether there is a single particle layer or, as is more often the case, a multiple layer or thin bed of material which is sufficiently dilated, permits the penetration of higher specific gravity grains. Thrust of Flowing Films. When a thin film of liquid flows over a plane solid surface, the layer next to the surface remains at rest but the velocity of the film increases with the distance from the surface and becomes a maximum near, but not quite at, the free surface. Therefore, a particle in suspension in such a film is acted upon by a greater force near the upper part of the film than at the lower part, resulting in an overturning effect. After a particle reaches the separating surface or an accumulated densely packed bed of other particles, the liquid flow causes it to move downstream by rolling, sliding, or by a movement involving alternating suspension and deposi¬tion (saltation). In rolling and sliding, which are brought about by a substantially noneddying stream, the large submerged particles are acted upon to the greatest extent and they move more rapidly than smaller ones, notwithstanding their greater mass. When two particles of the same size but of different specific gravity are considered, the higher density one moves more slowly by reason of its greater mass. As a result the particles tend to become arranged in the manner shown in Fig. 29. If any particle is so large as to stand above the water surface, the transporting force on such a particle, although the maximum available, will have less effect than that on a submerged one. Such large grains are therefore carried a lesser distance. Fluid Velocity in a Flowing Film It can be shown that if y is the distance within the flowing film from the interface, the fluid velocity v' at this point is given by IIn this equation p' is the fluid density, g the acceleration due to gravity, a the viscosity, a the angle made by the film to the horizontal, and 0 the film thickness (in cgs units). Furthermore, the volume Q of fluid per unit time and unit width is related to the film thickness and can be calculated, since which upon integration gives Consideration of the velocity v' (Eq. l) and the terminal falling velocity µ derived from Stokes law enables the horizontal travel dz in time dt to be calculated as follows: Again integration and substitution of Eq. 3 gives In this equation, z is the distance travelled before a particle of density p at the top of the fluid film has settled on the concentrating surface. From these expressions it can be seen that the depth of the flowing film varies as the cube root of the volume of fluid and inversely as the cube root of the sine of the plane surface inclination. Similarly the distance a particle can travel before reaching the surface is directly related to the quantity of fluid and its viscosity. The viscosity of pulp is greatly increased by the presence of near colloidal particles and consequently the penetration of the somewhat larger grains through the flowing film is retarded, resulting in them being carried further across the concentrator deck. Obviously this effect can be reduced by prior removal of some or all of the ultra¬fine particles. The detrimental effect of kaolin on the separation of fine ferrosilicon from quartz has also been demonstrated by Johnston,116 the relationship between viscosity and efficiency of separation being shown in Fig. 30. The pH value also has a very marked influence on the viscosity and the use of acid mine water, for example, can increase viscosity materially and thereby result in lower efficiency. Dilation of Multi-Particle Systems Except during washing (when a single particle layer may exist) concentration usually takes place in a bed of particles which is partially dilated or stirred by eddy currents. The extreme case is seen in various forms of the sluice where the flow of water and the presence of riffles combine to create a fluid bed in which reverse classification takes

Jan 1, 1985

By David Wilburn

Recycled materials are increasingly being used in road construction to supplement natural aggregate derived from crushed stone, sand, and gravel. An understanding of the amount of substitution that is taking place, its economics, and factors affecting the level of substitution is useful in estimating the potential for recycling and in assessing the total supply of aggregate. While U.S. consumption of recycled aggregates from crushed concrete is estimated to have increased 160 percent between 1994 and 1996; it comprised less than 5 percent of total aggregates consumed in 1995. Construction waste supply is regional, and is determined by infrastructure decay and replacement rates. Aggregate recycling rates are greater in urban areas where infrastructure is being replaced, natural aggregate resources are limited, disposal costs are high, or strict environmental regulations prevent disposal. Recycling is expected to increase as construction contractors attempt to save on transportation, disposal, and new material costs. Aggregate producers include recycled material as an option to consumers in order to prolong the life of their reserves and improve their range of products. Recycled aggregates can be a supplement for natural aggregates in selected applications, but their use should be evaluated locally based upon relative cost, quality, and market factors Because aggregate recycling is often affected by local and regional transportation systems, disposal and tipping fees, resource supply, the market for specific products, and municipal support, costs for three representative fixed site operations of different sizes were modeled. Under study conditions, all were found to be profitable, but highly dependent upon local tipping fees and market prices, which can differ significantly by location. Smaller operations were found to have different operational dynamics, often requiring creative marketing or incentives to maintain profitability.

Jan 1, 2001

By H. L. Ehrlich

Marine manganese nodules harbor a mixed population of manganese-oxidizing and -reducing bacteria, as well as bacteria that do not act on manganese. They also host phagotrophic protozoa. Out of several laboratory studies, one line of investigation indicates that some kinds of bacteria found on nodules can catalyze the oxidation of Mn2+ bound to Mn(IV) oxide at the surface of nodules at in situ temperature and pressure. In the absence of appropriate bacteria, the rate of oxidation of the surface-bound Mn2+ is significantly slower. How Fe(III) is incorporated is still unclear. Base metals such as Cu, Ni and Co, which are usually found in marine nodules in various quantities, may be incorporated through scavenging by Mn(IV) oxide. Laboratory evidence shows that some other bacteria on nodules can reduce the Mn(IV) oxide to Mn2+ in the presence of an appropriate source of reducing power in the form of organic carbon. The reduction of Mn(IV) oxide mobilizes manganese and trace metals such as Cu, Ni, and Co in the nodules. These bacteria reduce Mn(IV) whether oxygen is present or not. They could be harnessed in the bioleaching of nodules. The manganese(IV)-reducing bacteria and the bacteria that neither oxidize nor reduce manganese are thought to keep the organic matter in the nodule environment sufficiently low to favor bacterial manganese-oxidation, which in laboratory tests is inhibited by excess organic carbon. The preying by the protozoa on the different kinds of bacteria is thought to control the bacterial population on nodules and ensure continued growth of the manganese-oxidizing bacteria and the nodules. To eliminate interference from seawater salts in the bioleaching of Mn, Cu, Ni and Co, it may be more advantageous to use terrestrial Mn(IV)-reducing bacteria than the Mn(IV)-reducing bacteria resident on nodules. The terrestrial bacteria would allow bioleaching in fresh water. Unlike the marine bacteria, most terrestrial bacteria studied to date reduce Mn(IV) oxide only anaerobically. A general outline for an anaerobic bioleach process of marine nodules with freshwater bacteria is discussed.

Jan 1, 2000

The role of ZnSO4 in activation of sphalerite (ZnS) and deactivation of lead-activated sphalerite has been investigated. Adsorption of metallic ions at the solid-solution interface, electrokinetic results for sphalerite under different chemical conditions, and thermodynamic calculations were used to elucidate the mechanisms involved. It is shown that flotation response of sphalerite in absence of activators is rather weak and that adsorption of ethyl xanthate on the mineral mainly electrostatic in nature. Zinc sulfate is found to activate the mineral but only in a narrow acidic pH range. This is attributed to the increase of zeta potential as well as the possible increase in amount of negatively charged bulk precipitated zinc xanthate. Lead ions activate sphalerite in a wide pH range and the lead activated mineral assumes a flotation behavior similar to that of galena (PbS). However, addition of ZnSO4 depresses the lead-activated sphalerite by replacing lead ions at the mineral/solution interface.

Jan 1, 1999

By D. C. Torre

Ever since the introduction of the first successful continuous mining machines some twenty odd years ago, there has been an obvious, and steadily increasing, need for some form of continuous haulage machine, or system to work with them. There is universal agreement that the continuous mining principle, and the various continuous mining machines in particular, have done a great deal for mining. However, coal mining men, who have the longest and broadest experience with them, agree that continuous miners do not truly mine continuously. In fact, their estimates, and actual studies, show that the real mining time of these machines ranges, on the average, from 20% or less to about 30% of the available shift time. Part of the failure of the continuous mining machine to mine continuously may be ascribed to breakdowns in its intricate machinery. Most of the failure, however, is blamed on a variety of other factors. Of these, there are two which are given particular blame: face haulage and roof control. The relative importance of the two as the major negative factor in continuous mining is a moot point. However, there is an extensive concensus that haulage is the number one problem. Roof problems and their control are inhibiting factors regardless of the type of mining or mining machines employed. But the truth of the matter is that a continuous mining machine, like any type of loading machine, whether underground or on surface, is only capable of mining or producing in direct ratio to the mineable, or loadable, material that can be kept ahead of it and equally, if not more, important, that can be taken away from it. While the first commercially successful continuous miners were developed for, and proved successful in, coal mining, they were soon

Jan 1, 1974

By J. G. Clevenger

Introduction Computer-based haul truck dispatching has been operational at Tyrone since May, 1980. The system, known as Dispatch, has accomplished significant productivity gains compared with the former method of manual radio dispatching. These gains translate into a reduction in equipment requirements or an increase in production. An operational mine may choose to increase production or reduce the number of equipment shifts operated, resulting in substantial operating savings. Dispatch Description The core of Dispatch is a VAX 11/780 computer. Originally purchased with 512 kbytes of memory, the system is now upgraded to 3.75 Mbytes to accommodate users from all areas of the operation. Dispatch alone requires 200 kbytes of locked-in memory, with an additional 300 kbytes of memory available on a top priority basis. The computer, located in the Tyrone Branch general office building, is connected by under-ground cables to the dispatch tower, 3.2 km (2 miles) from the general office. Four two-paired lines are used to drive two cathode ray tubes, a printer, and the Dispatch master panel. One CRT continuously displays pertinent pit information for the dispatcher's use. Routinely displayed transactions include truck assignments, changes in shovel material type, truck and shovel breakdowns and delays, and changes in truck, shovel, or dump status. Er¬ror detection capabilities include late arrival and improper input from the field panels. The second display terminal is used for utility reporting and input functions. Status reports are displayed for availability, position, and material type being handled by any given truck or shovel. Fuel reports are available indicating the time last fueled, quantity of fuel delivered, and the time spent fueling. A truck factor utility reports relative performance to other trucks and a shovel utility reports efficiencies in tons per shovel shift. Instantaneous shift reports and mine status reports are available. Inputs such as lineups, operator changes, new travel times, shovel bias, tie down times, and lunch times are all handled with the utility display terminal. The printer records a hard copy of operating truck locations and production from the beginning of each shift. If there is a system malfunction, the printer copy is readily available to the foreman for radio dispatching. The Dispatch master panel is a microprocessor-based panel that allows the dispatcher to override any computer-assigned transaction. Classification of "bad order" conditions and acceptance of certain field messages is entered through the master panel. This panel communicates with all truck and shovel field panels via UH radio frequency through a repeater tower located on a high ridge contiguous to the mining area. Each truck and shovel field panel unit transmits a coded signal that is identified by the computer. A typical communication sequence between field panels and the computer for one complete truck cycle is: • The truck panel LED display indicates the shovel number to which it is assigned. The truck arrives at its assigned shovel and the operator presses the "Arrive" button on the truck panel. As the computer receives the "Arrive" message, it concludes the return (empty) travel time and begins accumulating idle time before loading. • When the first dipper of muck is placed into the bed of the truck, the operator presses the "Load" button. The computer concludes the truck idle time accumulation and begins the loading time calculation. • When the truck is loaded, the shovel operator presses the "OK"

Jan 9, 1983

By Leo Misagi

THE ROLE OF MANAGEMENT IN PLANT SAFETY Safe Production In plant management, safe production is an ideal. It becomes reality when: 1. Plant employees work in healthful and safe conditions. 2. Plant safety is in harmony with productivity. 3. Worker absenteeism and turnover are low-the work force is stable. Plant management must apply its talents, skills, and resources to build and maintain the foundations of safe production on a system of policies, procedures, and attitudes. It must do so by marshaling all managerial functions - staffing, direction, control, innovation, and representation-toward that goal. Ideally, a plant safety program is proactive. Management should plan and implement safety activities that are aimed at prevention. Past accidents-accomplished facts-should not be the sole guide of future actions. In other words, positive preventive measures, not the fewness of accidents are the criteria for safety production. An ideal safety program is one which is cost effective. Management should view plant safety as part of a broader program-loss control. Loss control includes prevention of fires and other accidents, unhealthful conditions, and occurrences that cause bodily injury, illness, property loss, or damage. Management evaluates the employees' productivity and safety. An effective training program aimed at loss control is the hallmark of safe production. Ideally, both labor and management should be well versed in safety. They believe in the premise that work-related injury and illness, as well as property loss and damage from fire and other accidents, are manageable and preventable. Such belief prompts them to learn more about the work process and potential hazards associated with each job. They also learn how to expect and prevent fues and other accidents, and respond to emergencies at the plant. Plant management determines which individuals or groups of employees are accountable for the following functions:

Jan 1, 1991

By Don Suttie, Frank Roybal, John Wright

The Roybal Raise incorporates the simplicity and low cost of a short bald-headed raise while overcoming the usual impediments to driving a high baldheaded raise. The concept provides for speedy excavation and maximizing the use of the excavated opening, using the rock for structural support. The design is simple: two baldheaded raises are driven parallel to one another, interconnected by dogholes about 8 m (25 ft) apart. In this fashion, the raise can be driven safely and efficiently to 60 m (200 ft) or more. When the raise is complete, the excavation requires no timber for support or partitioning. Rather, the rock itself supports the construction. As a test, a two-compartment Roybal Raise was attempted at the Independence mine in the Willow Creek District of Alaska. One side was driven 2 x 2 m (6 x 6 ft), intended for a manway and equipment skip. The other was driven 1.5 x 1.5 m (5 x 5 ft) for a muck compartment. Both sides were driven at a 65° angle from horizontal and 1 x 1 m (4 x 4 ft) dogholes spanned the 4.6-m (15-ft) pillar between the two compartments. After topping out at 50 m (160 ft), the raise was found to meet all expectations of efficiency, low cost, and practical simplicity. Technique Preparing the Roybal Raise for driving involves no elaborate or costly site prep except assembling materials and providing for muck removal. Normally, the raise muck falls directly onto the sill or track and is picked up either by an overshot mucker or a front-end loader. Raise mining proceeds initially by taking up both compartments simultaneously to the first doghole. It progresses by alternating one side as a clear manway for access with the other side, the active side, as a muck compartment (Fig. 1A). This stage proceeds quickly for no mucking or staging is required to cycle the first three rounds. On the fifth round, in addition to the normal raise round, rib holes are winged out toward the adjacent raise, thus starting the dogholes. Finally, as the sixth raise rounds are drilled and shot, the first doghole is also completed from both sides. The second stage commences by establishing one side as a clear manway for access while the other side, used as a muck compartment, is prepared for active mining. Figure 1B illustrates this step in the succession of raising with the active side advancing to the second doghole level. During this stage, services are maintained through the manway side and equipment is stowed in the doghole during blasting. As the active side advances, muck may be allowed to fill the bottom portion of that side since it is unnecessary to clear it for access. The next stage of raising involves cleaning out the former muck compartment and preparing it for use as a manway. At the same time, the former manway is prepared for advance and use as a muck compartment (Fig. 1C). As in the previous step, the same doghole is used for access and equipment storage. The active side advances until the second doghole is completed, at which point equipment and access are moved to the new doghole. From there, the active side is driven past the second doghole 8 m (25 ft) more to the third doghole level. Again, half of the doghole and the next raise round are mined. The fourth stage involves switch ing back the manway side for the muck side (Fig. ID). The new muck side is driven to the third doghole where equipment storage and access are reestablished. From there, the active side is advanced to the fourth doghole. In the same fashion, the raise is advanced to the top, alternating the manway side with the active side, maintaining access through the inactive manway, and storing equipment in the uppermost doghole. The cycle time for driving each compartment, regardless of the raise height, is greatly improved over that for normal bald-headed raises of comparable height. By storing equipment in the upper-

Jan 1, 1984

By W. Jack Wilkes, Gilbert L. Butler

TRANSPORTATION TUNNELING The use of tunnels for transportation, namely rapid rail transit systems and highways is on the increase. How much that increase might be through the 1980's is the question that is discussed in this paper. In 1975, there were 218 route miles of subway tunnels for rapid rail transit and 154 lane-miles of highway tunnel. From the viewpoint of many transportation planners and designers, the tunnel is one of the tools which may be used to by-pass many types of obstacles: a mountain, a body of water, or a Central Business District (CBD). Therefore, the transportation tunnel is a rural as well as an urban solution with major emphasis on the latter because of the current transportation needs of our cities. In the case of rural locations, tunnels are usually constructed through mountains in order to eliminate miles of steep grades and dangerous curves, such as the Eisenhower Tunnel which cuts through the Continental Divide in Colorado along 1-70 and saves 10 miles and 40 minutes of driving time. Tunnels also provide an excellent transportation link across bodies of water in order not to interfere with shipping lanes. The Chesapeake Bay Tunnel and Bay Tubes of the BART system are excellent examples of this type of tunnel.

Jan 1, 1976

By Kenneth A. Gutschick, Harry L. Francis

A large number of minerals and chemicals (Table 1), some mineral-derived, are used in the treatment of both potable water and domestic sewage waste water and solids (Anon., 1990). In water treatment there are ten principal process uses as shown in Table 2 (Anon., 1990), which also lists the major chemicals in each process. Since lime (calcium oxide, calcium hydroxide, and dolomitic lime) is the largest chemical (by tonnage) used in water treatment, this chapter will deal primarily with this product. Note that it is included in five of the process uses shown in Table 2, and that indirectly it is also used in at least one other category (taste and odor control, where lime is used with potassium permanganate to provide optimum pH condition). POTABLE WATER TREATMENT Water Softening The biggest use for lime in water treatment involves water softening, where it serves to reduce carbonate hardness caused by dissolved bicarbonate and carbonate compounds. Currently there are more than 1000 domestic water softening plants (Table 3) located in most sections of the country, but particularly in Ohio, other midwestern states, Florida, and Texas, where limestone bedrock abounds (Anon., 1990). Among the other uses for lime in water treatment are coagulation, disinfection, and removal of heavy metals due to high pH softening, removal of certain radioactive and organic compounds and fluorides, neutralization of acidic water, removal of silica, particularly with dolomitic lime, and removal of color. Most of the above advantages accompany water softening, although at some plants lime is used specifically for one purpose, such as silica removal with dolomitic lime, minimizing pipeline corrosion by addition of lime to correct an aggressive water, and others. Hardness in water comprises two types - carbonate hardness, caused by dissolved calcium and magnesium carbonates and bi- carbonates, and non-carbonate hardness caused by dissolved non- carbonate salts, particularly calcium sulfate, calcium chloride, magnesium sulfate, and magnesium chloride. Total hardness is the sum of the carbonate hardness and non-carbonate hardness. Waters are classified as soft (0 to 75 mg/L as CaCO,), moderate (75 to 150 mg/L), hard (150 to 300 mg/L), and very hard (above 300 mg/L). Most water plants strive to reduce total hardness to about 100 mg/L, although in recent years the tendency in some cities has been to allow a little higher value in order to reduce both chemical costs and the amount of water softening sludges produced. There are four basic types of softening plants: lime softening, primarily to reduce carbonate hardness; lime-soda softening, for reduction of both types of hardness (in some plants caustic soda is used in place of soda ash, but it is more costly); split lime-zeolite, where a base exchange mineral is employed in place of soda ash (this system also requires salt [sodium chloride] to regenerate the zeolite mineral); and finally zeolite softening, which is generally used in small systems for private houses and small municipalities. Principal disadvantages of the latter system are its relatively high cost and the salt requirement, which introduces considerable quantities of soluble salts into the waste water discharge system.

Jan 1, 1994

By G. Chi, M. C. Fuerstenau, R. C. Bradt, A. Ghosh

Technical Papers, Mining Engineering Vol. 49, No. 12, pp. 70-75, Discussion by McIvor, R. and Fink, R. P. We must challenge what we consider to be dubious statements in the above-mentioned article. Much of what is reported was published elsewhere over a year ago (i.e., in their reference: Chi, G., Fuerstenau, M.C., Brandt, R.C. and Ghosh, A., 1996, "Improved comminution efficiency through controlled blasting during mining," International Journal of Min¬eral Processing, Vol. 47, p. 93). There is nothing unusual or significant about a 10% or more difference in grindability between two materials of apparently similar origin and other attributes (such as chemical assays, hardness and specific gravity). Consider also that: • the materials came from different locations in the pit; • only two sets of four blastholes were considered per blast: • spatial orientation with respect to the face and other blastholes was also altered; • only two pairs of data were collected; • the impact of sampling error and bias, • both of which could be significant in this case; and • normal experimental error in parameter measurement and calculation. It is clearly inappropriate to draw any conclusions with regard to cause-and-effect relationships without first presenting a thorough error analysis. A substantial number of randomly selected test and control areas might provide meaningful statistical evidence of the claimed effect of the variables being studied. Even that appears doubtful, however, because Figures 3 and 4 show about as much variation in crushing energy for the two "regular" conditions vs. regular and altered conditions. On a separate, but also important issue, broad experience with autogenous/semiautogenous milling has shown that smaller or weaker feed material results in reduced tonnage. ? Reply Thank you for the opportunity to respond to the comments of Messrs. Mclvor ant Fink. The points raised are addressed in the order of presentation. As the commenters noted, there are parallels between this article and a previous manuscript. The reason for these complementary articles is that the audiences for the two publications are different, although there is obviously some overlap. We believe that the different audiences each profit from the technical aspects of the two manuscripts. We are not in agreement that a grindability difference of 10% is insignificant. The laboratory work is corroborated by data obtained at plant scale, which confirms that an increase in throughput of the same magnitude (between 5% and 10%) is achieved when the blasting pattern is modified, i.e., reduced from 9 to 8 m (29.5 to 26.2 ft). This increase in throughput translates into substantial power savings for ore grinding. No differences in appearance in the rock were noted between the two adjacent areas. Rock in both areas belonged to the same rock mass/block, which was quite homogeneous and without geological features in between that could create drastic differences in comminution characteristics. Given the resources, we would have sampled from all of the holes. Additional data would certainly have been desirable, as would have been additional funding to expand the study. Their comment, "spatial orientation with respect to the face and other blastholes was also altered," is not clear. Figure 1 does not show the initiation sequence. Without the initiation sequence, we do not understand how the commenters concluded that spatial orientation with respect to the face and other blastholes was altered. The commenters may be trying to make the point that the four-hole configurations with respect to the original bench face are not exactly the same. In fragmenting the

Jan 1, 1999

By D. A. Elgillani, N. A. Abdel-Khalek

The role of ZnS04 in activation of sphalerite (ZnS) and deactivation of lead-activated sphalerite has been investigated. Adsorption of metallic ions at the solid-solution interface, electro kinetic results for sphalerite under different chemical conditions, and thermodynamic calculations were used to elucidate the mechanisms involved. It is shown that flotation response of sphalerite in absence of activators is rather weak and that adsorption of ethyl xanthate on the mineral mainly electrostatic in nature. Zinc sulfate is found to activate the mineral but only in a narrow acidic pH range. This is attributed to the increase of zeta potential as well as the possible increase in amount of negatively charged bulk precipitated zinc xanthate. Lead ions activate sphalerite in a wide pH range and the lead activated mineral assumes a flotation behavior similar to that of galena (PbS). However, addition of ZnS04 depresses the lead-activated sphalerite by replacing lead ions at the mineral/solution interface.

Jan 1, 1999