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By Harold D. Taylor

Cone fitting is a procedure which can be used to estimate the short range, linear portion of a variogram. Rather than combining data over windows In different directions and fitting by hand, cone fitting estimates the variogram in all directions at once using a multiple regression approach. The method automatically provides estimates of nugget effect, anisotropy, and initial slope of the variogram. This paper presents the results of simulation studies which have been carried out. These studies Indicate that the method works well and that it may be particularly useful for small data sets. The simulation studies also indicate that the local variogram may differ greatly from the theoretical variogram.

Jan 1, 1986

By D. U. Deere, T. R. Harrell

Factors such as rock cover, rock type and rock quality can exert a dominate influence on the behavior of a concrete tunnel lining. This paper will present and discuss the investigation, evaluation and use of such factors at the Rocky Mountain Pumped Storage Project in the design of the 10.67m (35 foot) diameter concrete tunnel lining. A plane strain, finite element analysis was performed on the concrete/rock system. The finite element model consisted of a concrete lining zone and three (3) rock zones; a highly destressed rock zone, a transitional rock zone and an intact rock zone. Zones were differentiated from one another by varying the modulus for each zone. Results of this study yielded a concrete lining strain from radial expansion which was used in the crack control calculations for the concrete tunnel lining. After the tunnel excavation was performed, plate jacking tests and additional seismic surveys were run on the tunnel walls. This testing served to confirm the rock modulus values used in the preliminary analysis. However, results were observed which indicated that a two zone rock model would more accurately represent in-situ rock conditions than the three-zone rock model used in the preliminary analysis. This revision to the concrete/rock model was made and a final concrete tunnel lining analysis was performed.

Jan 1, 1986

By E. Topuz

The objective of mine ventilation is to provide suitable environmental conditions in working places at an economical cost. Maintaining a stable air flow, i.e., a flow quantity and the direction, is of vital importance. In practice, however, air flow quantity in underground openings varies continuously. Therefore, the stability of air flow means relative stability that is the flow in a branch moves along a certain direction while the variation in air quantity stays within acceptable limits. Many underground mines have low stability of mine air flow because almost all types of ventilation systems involve complex diagonal branches through which air flow quantity and, more importantly, the air flow direction may change easily. This creates extremely hazardous, unstable conditions, especially during emergencies. This paper deals with the directional stability of air flows in underground mines in single or multiple fan systems. It examines the rules of directional changes of air flows in diagonal branches, and proposes a method of calculating the stability conditions. It provides a criterion that may be used in designing new mine ventilation systems to ensure a desired degree of stability or in improving the existing systems.

Jan 1, 1986

By D. Wheeler

Similar to modern day petro-chemical processes, the flotation column is a complete break from the conventional manner of floating on a batch basis in cells. The column has a reputation of being only suited for cleaning applications, particularly that of molybdenum cleaning. The original column, invented and developed by Column Flotation Co. of Canada, Ltd., was developed as a Rougher-Scavenger. Only after proving its capability in this field was cleaning allowed. One rougher-scavenger column in closed circuit with a cleaner column produced better grade and recovery than a conventional cell plant consisting of one stage of roughing, two scavenging and four stages of cell cleaning. Three cleaner columns replaced 13 stages of conventional cell cleaning and resulted in higher grades and recoveries. The gentle countercurrent action of the column itself together with complete automatic control enable this machine to produce such results.

Jan 1, 1986

By Herschel H. Potter

While preparing this paper, I soon realized I had a problem in defining what injuries should be classified as "personal." Admittedly, in some cases, falls of persons, sprains and strains, pulling and pushing injuries, and back injuries may have been caused by the individual. But, I believe the number of injuries directly attributed to an individual's carelessness is relatively small in comparison to the total number of injuries in those categories. For discussion purposes, the injuries most often considered to be personal, that is, caused by individuals are those classified as back injuries. My paper will be devoted to back injuries because this category represents the greatest number of injuries occurring in coal mining. A man's best friends are his wife, his back, and his dog. The back, however, has a reputation of not being as faithful as his other two friends.1 "Industrial back injury is one of the major occupational health problems plaguing industry to- day. According to national surveys, back injury accounts for more lost time than any other occupational injury and costs companies thousands of dollars in compensation, medical fees and over- time pay."2 A study conducted by USA Today found that back ailments afflict an estimated 75 million Americans, have totally disabled about 2.5 million adults, lead to as many as 200,000 required surgeries each year, are a leading cause of absenteeism and disability, are the second leading cause of hospitalization, are responsible for 18 million doctor visits a year and cost about 93 million work-days annually. The cost to industry is estimated at $14 billion a year.3 During the past 30 years, a great deal of accident prevention education has been directed toward the reduction of back injuries. However, this effort has not produced any great change in the incidence of the injury. Back injuries should be of the utmost concern to everyone here today. I don't need to tell you what the incidence rate is for back injuries, nor will I use any statistics to inform you of the incidence rate for the coal mine industry or any specific mine. It is only important to know that the number of back injuries is excessive. Back in- juries resulting from lifting have been and continue to be one of the most constant and undiminishing problems in the occupational environment. The fact is, we use our backs to per- form work that the back was not designed to do. Historically, manual materials handling related accidents have consistently accounted for the largest percentage of all industrial accidents. In fact, over 25 percent of all industrial accidents are related to handling of materials.4 In underground coal mining operations, material handling accidents account for even larger percentage of lost time injuries. In 1983, for example, they accounted for 34 percent of all lost time injuries and 32.6 percent of all non-lost time injuries. 5 I want to use one example of a task that will illustrate what I have just said: the building of a concrete block stopping. Each solid concrete block will weigh from 60 to 70 pounds. To build a stop- ping 6 feet high and 20 feet wide will take approximately 80 blocks. On the average, these 80 blocks will be handled at least four times from loading into the conveyance for transportation to being placed while constructing the stopping. This generally requires repetitive lifting from ground level. The next three slides demonstrate the lifts

Jan 1, 1986

By Yoshiak Mizuta, Hi Keu Lee

The manuscript of this paper was not available for publication when the Proceedings went to press. Should the authors elect, they may distribute copies at the time of the Symposium

Jan 1, 1986

By Someswar Kesh, Gavin J. Faulkner

Introduction Powered supports in longwall mines essentially operate under two conditions: 1. Static (non-working) 2. Dynamic (working) There is a variety of static conditions which may exist in a mine. These may include a temporary stoppage of work for a few minutes or stoppage of work for days because of strikes or lack of demand for coal. Studies on pressure behavior were made during brief stoppages of work (few hours) as well as a long stoppage of work (two weeks). In dynamic conditions, the pressure variation on supports is expected to follow a certain pattern. However, in many cases that does not happen. Pressure-time characteristics of each leg of the supports were studied. This paper discusses the research findings of the studies made on powered support pressure behavior under both static and dynamic conditions. Studies on Pressure Variation in Static Conditions It has been noted (Brenner et a1.,1975) that the pressure distribution on mechanized longwall faces is random in nature and changes with time and position along the face. Figure 1 and Figure 2 show the distribution of support pressures for the front left leg of supports in two different mines. In both the cases it can be seen that the variation of the support pressures along the face is indeed absolutely random. However, the nature of randomness in both the cases is very different. The data in Figure 1 (Mine 1) is much more random than the data in Figure 2 (Mine 2). This can also be seen from the standard deviation of the data (81 26 Pa versus 4000 pa) . By comparing the results of ten other sets of data the same conclusions can be arrived at. Since it was found that in Mine 1, the pressure on support legs was extremely random, batches of ten supports each were taken and the mean pressure plotted. This mean pressure line is shown in Figure 1. It was expected that this would smooth out the variation due to individual props and reveal a better picture of the support pressure distribution. For the case cited, the standard deviation is lowered to 5043 Pa from 8126 Pa which indicates that, on an average, the supports tend to take loads more uniformly, however for individual support legs the load on the supports is very random. The significance of this is that, if the supports are selected on the basis of theoretical load calculations alone (based on the roof block concept or some other concept), only the mean value of the support load required will be known. In the case of a pressure distribution with a high standard deviation, some supports selected on the basis of mean value will fail, while others will be supporting low loads. Hence, for proper selection of support capacity, it is not only necessary to know the average load but also the expected nature of the distribution. The only reason why the supports are supporting the roof in Mine 1 well is the fact that the yield load is above the highest level of load attained. If Mine 2, which has a much higher average load, had the same sort of standard deviation, the loads on the individual legs could easily exceed the support load. For example, if the pressure distribution followed a normal curve and the mean pressure was 36,700 Pa with a standard deviation of 9000 Pa, it is possible that 67% of the supports will be within the pressure range of 36,700 plus or minus 9,000 Pa, or 22,700 and 45,700 Pa. Many of the supports could then exceed the yield load of 40,000 Pa. It is clear then that the mean load is an inadequate criterion for estimating support resistance. The nature of the problem is statistical and such an analysis should therefore be made as far as pressure data are concerned. Statistical Analysis of Pressure Data. In general, in longwall mining and other geo-mechanics studies, it is extremely difficult to specify the exact nature of the distribution of a variable. Under these circumstances, probability

Jan 1, 1986

By V. R. Degner

Given: the process tonnage, performance objectives, and plant floor area, the paper will outline the metallurgical and machine hydrodynamic scale-up procedures for translating the lab flotation test results to the design of a full size flotation machine installation, including the machine number, size, and arrangement. Machine selection considerations for rougher, scavenger and cleaner duty, and special considerations relative to specific applications, will be summarized where appropriate. The value of kinetic modeling in quantitative metallurgical (lab to full size machine) scale-up will be reviewed considering a coal installation, and design information for additional applications such as copper, molybdenum, iron, phosphate, and potash will be discussed.

Jan 1, 1986

By Bernard J. Reddy

This paper describes in detail the economic forces affecting cobalt supply, demand, and prices. The major factors that will determine the growth or decline in the use of cobalt in each of its applications are discussed. The flows of cobalt from mine to refinery are described on a company-by-company basis. The paper concludes with a discussion of how the characteristics of the supply and demand sides of market interact to determine cobalt prices.

Jan 1, 1986

By Bruno C. Boik

INTRODUCTION Cost Management is the evaluation and control of project scope, budgets, and costs with the objective to minimize costs and maximize the benefits to be derived from the use of project resources in conformance with Criteria stipulated or approved by the Owner. In order of importance, Cost Management is one of the most prominent Project Management functions because of its impact on the financial success of the venture. Cost Control and Cost Accounting are essential parts of Cost Management. The principal objectives of Cost Control are to monitor project commitments and expenditures, and measure performance for the detection of variances from acceptable standards and from control budgets so that timely action can be initiated to minimize costs. Cost Control is evaluation of current costs and projection of future costs, rather than the reporting of actual costs. Cost Control is not Cost Accounting. The primary objectives of Cost Accounting are the determination, classification, analysis and reporting in useful detail of actual costs. Following are some considerations on Cost Management and Cost Control as practiced by Owners and Contractors of industrial projects, Mining and Metallurgical projects in particular. COST MANAGEMENT Functions Conforming with the requirements of Owner's Criteria, the principal Cost Management functions are to • evaluate estimates of project costs and establish budgets for a scope of work and schedule; conversely, evaluate the scope of work that can be executed most effectively within a budget and schedule. • promote investigations and comparative evaluations of alternate designs and plans to find the most cost effective way to execute a project and minimize its operating costs. • based on Cost Control and Cost Accounting input, evaluate and initiate appropriate actions to correct deviations from approved budgets. assess project economics, the effect of changes on project economics and, if necessary, revise Project Criteria to modify their impact on costs.

Jan 1, 1986

By John Matthews

During 1982 and 1983, it became increasingly clear that North River No. 1 Mine would have to buy a set of replacement longwall shields. Despite extensive efforts to make improvements, the shields that were in place at that time were incapable of holding the roof. North River Energy had limited experience in specifying and acquiring longwall roof supports, but a great deal of experience with our conditions and needs. We there- fore compiled a shield evaluation method that uniquely responded to our situation. The resulting system emphasized teamwork among operating, maintenance and engineering personnel. The charge from management was threefold: select shields to support maximum expected loads, optimize face operations, and pay attention to detail. Final selection was made objectively, based on all available information. We chose a 5516 KN (620 tons) support with a minimum yield load density of 760 KN/m2 (7.9 tons/ft2). This paper reviews our process of specification, design, prototype testing, and fabrication, and offers some relevant suggestions.

Jan 1, 1986

By V. Nemec

The paper deals with further progress of the space and time models achieved since some of the previous APCOM symposia (1970, 1975, 1980) where original methodology was presented. Various transformations, de- and retectonization of space models, transfer of patterns, interactive time models are among the most important improvements. Enlarged possibilities of practical applications of the developed system are shown.

Jan 1, 1986

By T. S. Golosinski

This paper identifies the components of belt conveyor shifting force as a vertical force. Pv, horizontal force, Ph, and longitudinal force, P1. A method for estimation of these forces is proposed, which relates their value to the properties of the conveyor structure and rails, shifting operation parameters, and other site-specific conditions. The effect of this force on shifter behaviour is analysed and guidelines are developed for selection of a suitable shifter. Also discussed are the requirements of the shifting operation for other equipment.

Jan 1, 1986

By Christopher J. Bise

INTRODUCTION Traditionally, the purpose of mine ventilation has been defined as diluting, rendering harmless, and sweeping away hazardous gases which may accumulate in the mine workings. In addition, the ventilation sys- tem must keep the mine air as close as possible to the composition of the outside air in order to provide a healthy environment for mineworkers and to prolong the useful life of mine equipment. When plans are being developed for opening a new mine, sound engineering dictates that a calculated estimate be made of both the maximum air quantity requirement and the accept- able mine pressure (head) loss related to the under- ground layout and projections. For efficient mine ventilation, the two problem areas that demand constant attention are (1) the control of fugitive air (through short circuiting or leakage) and (2) the provision of ample airway capacity. Fugitive air is that quantity of air which enters the mine but short-circuits the mine ventilation system through improperly constructed stoppings, doors, overcasts, etc.; it represents an investment in power consumption (to move the air by the fan) while providing little, if any, return on the investment. Typically, efficient mines are those that can deliver 50% of the air handled by their fans to the working faces. Paying close attention to a new mine's ventilation system at the outset can provide numerous benefits during the later stages of mine development. It has often been noted that the main cause of trouble in many mine ventilation systems can be found within a radius of 2500 ft from the fan installation. Ample airway capacity is essential to maintain the mine's head loss within manageable limits. Similar to an industrial ventilation setting, the mine's headings are the ductwork for the handling of air. The difference between normal industrial ventilation and mine ventilation is apparent when more capacity is needed; a plant can install new or larger ductwork, but it is difficult for a mine to install new entries or enlarge existing entries. Air velocities are usually maintained under 800 fpm to limit the stirring up of dust in the intakes. With this in mind, it is usually simple to estimate the number of entries required in the ventilation system. The problem that often arises, though, is when the system's capacity is overestimated and the required head loss to achieve the desired quantity is excessive. Experience has shown that problems are often encountered when the mine's condition requires much more than three inches of water gage to ventilate the mine on a day-to-day basis. Thus, if proper attention is not paid to these parameters in the preplanning and initial development of a new mine, it may take an extremely high water gage to achieve the ultimate quantity. This chapter will not delve deeply into the theory of mine ventilation; many other books have been writ- ten on that subject. Instead, the following sections will point out the important considerations that should take place before the development of a mine ventilation system, as they relate to normal industry practice. FUNDAMENTALS OF AIRFLOW Volume Versus Water Gage The common method of measuring ventilation pressure is in terms of inches of water gage (wg); one inch of water corresponds to a pressure of 5.2 psf (pounds per square foot). Further, airflow follows a square-law relationship between volumes and pressures where, as shown in the previous section, a doubling of quantity requires a quadrupling of pressure. Table 1 shows how heads and quantities are related. Using Table 1, it is readily apparent that, for the same mining conditions, an increase in pressure from one inch wg to three inches wg results in only 34,640 cfm when the initial quantity is 20,000 cfm. Pressure Losses Pressure losses in mine ventilation systems are caused by friction between the heading walls and the air stream and by shock losses due to abrupt changes of the air stream caused by obstructions. The common pressure-loss formula for mine ventilation is: [ ] where H is the pressure loss in inches of water gage, K is the coefficient of friction, L is the length of the airway in feet, 0 is the airway perimeter in feet, V is the air velocity in feet per minute, and A is the cross-sectional area of the airway in square feet. Table 2 lists the US Bureau of Mines schedule of friction factors (K) for mine airways. Recent research indicates that the values of K are slightly less for coal mines than those presented in the USBM Schedule, and Table 3 should be consulted in this instance. It should be

Jan 1, 1986

By B. Hagemeyer

The trend to higher solids paper coating and to neutral/alkaline papermaking is causing an above average growth in the use of calcium carbonate as both a paper coating and a paper filling pigment. These developments are expected to continue well into the 1990s. Thus, calcium carbonate will account for a somewhat larger percentage of total paper pigment volume. Nevertheless, overall growth in paper tonnage coupled with an increase in filler pigment percentages will create a demand for larger quantities of other pigments as well. Kaolin will continue to be the dominant paper pigment by far, with continued growth though at a somewhat slower rate than calcium carbonate.

Jan 1, 1986

By Otis M. Clarke

Observations of the land surface show that subsidence over active and abandoned underground mines is controlled to some extent by the stratigraphic and structural geology of the mines in addition to the methods of mining and the thickness of the seam mined. Exposures in highwalls of surface mines that intersect abandoned underground mines in the Warrior coal basin of Alabama illustrate this, showing sandstone beds that act as beams bridging openings and interbedded shale beds crumbling. Stratigraphic changes observed in the district are extremely variable and calculations regarding subsidence used for one area will not apply to another area if there is a major change in rock types Rocks exposed in road cuts and open pit mines demonstrate a marked difference in weathering characteristics. Some rocks, such as shale, wea- ther very rapidly, by crumbling and becoming plastic, whereas other rocks, such as sandstone, resist weathering. The continuing rock weathering after mining in openings produced by the mining activity and in fractures caused by roof movement, apparently, is a major cause of the surface sub- sidence that may occur years after the mines are abandoned.

Jan 1, 1986

By K. F. Unrug

The identification of the major subsidence areas in the Western Kentucky Coal Field has been made to distinguish the subsidence phenomena on the basis of geographical location Further, this identification process is based upon selected critical geological, topographical, and mining features in the coal fields. The proposed identification of the regions is derived from a comparison of the subsidence potential in a particular area. The most likely severe subsidence areas are compared to those with the probable low subsidence severity. The combination of certain geological and mining parameters involved allows for the derivation of a classification of the subsidence potential. It is recommended that further site specific studies should be performed on this area.

Jan 1, 1986

By A. W. Khair

This paper presents an analysis of surface subsidence characteristics in room and pillar mining using physical models and laser holographic interferometry (holometry). The analysis included the effect of various geometric parameters and different overburden materials (sand and gelatin) and resulted in the formulation of a more realistic model material for laboratory simulation of typical geologic overburden. Previous work using this technique involved the use of sand to model over- burden and detailed results are presented elsewhere (1). The main emphasis of this work was to develop a gelatin based mixture which would provide a better model. Initial work during the development of this model included testing various ingredients and monitoring the time dependent characteristics of the model, such as the surface penetration resistance. This value indicated changing deformability/ridgidity of the model with time. Once several models had been developed, holometry techniques were applied. Initial tests showed that the stiffness of gelatin created a bridging effect over the mine models, thus restricting deformation. At this point, a 0.5 inch (1.27 cm.) layer of sand was placed directly over the mine model with the gelatin mixture on top. The approach was to model different zones of overburden movement that exist as an extraction occurs (2). The sand would represent the caved and fractured zones which occur above the excavated area. The gelatin models the continuously deformed zone which occurs between the fractured zone and the surface. Holometry techniques applied to this model yielded successful results.

Jan 1, 1986

By Chun K. Ip, Simon T. Johnson, Robert J. Fowell

Laboratory conducted cutting tests at traverse speeds up to 1 m/s have been undertaken with high pressure water jets assisting the cutting action of drag tools. Four sandstones and one limestone have been investigated, which represents a wide range of compressive strength. Variables investigated include water jet pressure, nozzle diameter, depth of cut, traverse speed, jet position, and tool bluntness. The test programme has allowed a model to be developed which describes the action of water jet assistance for three conditions dependent on the ability of the jet to penetrate the rock: where no significant penetration occurs, where penetration is greater than 50% of the tool depth, and an optimum condition around 30% of tool depth.

Jan 1, 1986

By Dale S. Preece

INTRODUCTION The Waste Isolation Pilot Plant (WIPP) is a research facility located in Southeastern New Mexico. The WIPP is being developed by the U. S. Department of Energy (DOE) to demonstrate the safe storage of defense-related radioactive wastes in bedded salt. Since the WIPP is a research facility. a number of large-scale in situ experiments have been planned and are currently under construction (Munson, 1983). These experiments were designed to study various aspects of nuclear waste storage in bedded salt such as mechanical responses and creep closures of drifts, thermal response due to heated canisters, and thermally-induced fluid migration. One purpose of the WIPP is to develop and demonstrate a general predictive structural analysis capability for a bedded salt repository. One set of experiments, called the Geomechanical Evaluation (Room G), is heavily instrumented to study the creep around several rooms with different configuration. A Layout of the Geomechanical Evaluation room is shown in Figure 1. One portion of it consists of wedge shaped pillar where two drifts intersect at an angle of 7.5 degrees. The pillar can be seen at the bottom of Figure 1. One purpose of the wedge pillar experiment is to study progressive pillar failure in the tip region. Another is to determine how nonuniform pillar thickness affects creep closure of the drifts. An important aspect of the experiments is the correlation between experimental data and corresponding pretest finite element analyses of the site. The finite element simulations serve two purposes. First, the computer simulations aid in understanding the experimental results by providing calculated stress. strain and displacement fields that cannot be measured directly. Second, comparison of calculated and measured drift closures serves as a method for validating or improving the finite element models and the constitutive models employed. The wedge pillar geometry required a 3-D finite element creep calculation. Results from these 3-D calculations will be presented in this paper. A 2-D plane strain double drift model which approximates the wedge pillar geometry at a slice perpendicular to the drift has also been performed and will be compared to the 3-D results in this paper. FINITE ELEMENT COMPUTER PROGRAMS Two finite element computer programs. JAC (Biffle, 1984) and JAC3D (Biffle. 1986), were used. JAC is a 2-D finite element program developed for quasi-static analysis of non-linear solids. It employs the conjugate gradient iterative technique to obtain a solution Spatial integration is performed using a single gauss point in each four node quadrilateral element. An hourglass viscosity technique is used to control the zero energy modes that occur with single point integration. The single point integration combined with the explicit nature of the program and exploitation of CRAY-1 computer architecture results in very efficient execution J4: results were compared to results obtained with eight other 2-D structural creep computer codes in the second WIPP benchmark exercise (Morgan, 1981) JAC3D was derived from JAC to treat 3-D finite element models and has many of the same characteristics including single point integration and hourglass viscosity. These characteristics have made 3-D creep analyses more reasonable by significantly reducing computation time The first exercise of the creep capability

Jan 1, 1986