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By E. D. Gardner, William O. Vandenburg

This chapter on the square-set system of mining em¬bodies the results of a special study of the subject made as a part of the US Bureau of Mines (USBM) general investigation of mining methods and costs. Data for the chapter were gathered by the authors during visits to mines using square-set mining, from information circu¬lars prepared by consulting engineers to the USBM on mining methods at individual mines, and from articles in the technical press and the Transactions of the American Institute of Mining and Metallurgical Engineers. Al¬though some of the properties discussed are no longer in operation, the methods described are still valid. The square-set system of mining has been very widely applied and is one of the oldest methods in use in the US. It is mainly employed for extracting rela¬tively high-grade ore in heavy ground. It is also useful in taking out pillars left in mining by other methods. It is used to some extent as an auxiliary method with many underground systems and as an alternate with cut-and¬fill, shrinkage, and open-stope methods when heavy ground is encountered. In some cases as much as 50% of the ore may be extracted by square setting at mines listed as using some other method of mining. A large tonnage has been mined yearly by square setting, prin¬cipally in the western states. Due to the high cost of timber and the relatively low average production per manshift, this system generally is not used when the ore can be mined by some other method; it usually is adopted as a last resort. Square setting, however, has a special field in that ore bodies not minable by other methods can be stoped by this system. As square-set mining was widely known, there was a tendency for years to use the method regardless of its suitability. With the increasing need for economy and the development of other methods, square setting has been superseded in many mines by cut-and-fill, top slic¬ing, or sublevel caving methods. The term square-set mining as used in this chapter includes all stoping methods in which the space made by the removal of the ore is timbered with framed sets. The sets usually are square in horizontal section. The stope may or may not be filled with waste rock as stoping progresses. HISTORY OF THE SQUARE-SET SYSTEM Prior to 1860, open-stope methods with stulls for casual support accounted for most of the mineral pro¬duction from underground mines. Attempts to mine large ore bodies that required artificial support resulted in disastrous cave-ins and loss of life. The art of mining received a tremendous impetus following the introduc¬tion of methods for mining the bonanza ore body which was discovered in 1860 in the Ophir mine on the Comstock lode, Virginia City, NV. This ore body was phe¬nomenally rich and eventually proved to be 122 m (400 ft) high, 27 m (90 ft) wide, and 98 m (320 ft) long. The mining of this large high-grade ore body taxed the best engineering talent of the time; the problem was finally solved by Philip Deidesheimer. According to Eliot Lord (1882): "At the 50-foot level the vein of black sulphurets was only 3 or 4 feet thick, and could be readily extracted through a drift along its line, propping up the walls and roof when necessary by simple uprights and cap-posts. As the ledge descended the sulphuret vein grew broader until, at the depth of 175 feet, it was 65 feet in width, and the miners were at a loss how to proceed, for the ore was so soft and crumbling that pillars could not be used to support the roof as in coal mines. They spliced timbers together to hold up the caving ground; but these jointed props were too weak and ill-supported to with¬stand the pressure upon them, and were constantly broken and thrown out of place. The dilemma was a curious one. Surrounded by riches, they were yet un¬able to carry them off and their mass of black sulphurets bade fair to become a white elephant on their hands. The Ophir Company began to wish themselves less for¬tunate, as their miners narrowly escaped burial day after day in their attempts to stope out the ore. "A young mining engineer, Philip Deidesheimer, was in charge of a quartz mine in Georgetown, El Dorado County, Calif., in the autumn of 1860, when this serious check to the development of the lode occurred. At the request of William F. Babcock, a trustee of the Ophir Company, Mr. Deidesheimer left his California mine and crossed the mountains with a letter from the direc¬tors of the Ophir Company, authorizing him to inspect the workings of their mines and make such changes in the method of timbering as should seem to him expe¬dient. After examining the vein he designed, in the course of a few weeks, a system of timbering which proved to be exactly adapted to the requirements of the work. Experiments which he had previously made in California gravel and quartz mines furnished the out¬line of his plan. This was to frame timbers together in rectangular sets, each set being composed of a square base, placed horizontally, formed of four timbers, sills, and crosspieces from 4 to 6 feet long, surmounted at the corners by four posts from 6 to 7 feet high, and capped by a framework similar to the base. The cap pieces forming the top of any set were at the same time the sills or base of the next set above. These sets could readily be extended to any required height and over any given area, forming a series of horizontal floors, built up from the bottom sets like the successive stories of a house. The spaces between the timbers were filled with waste rock or with wooden braces, forming a solid cube when¬ever the maximum degree of firmness was desired. By adjustments of these sets the ore bodies along the line of the lode were extracted with comparative ease and security." Most mining methods are evolved slowly and are the result of the combined efforts of a number of minds; Deidesheimer's feat was remarkable in that his system was immediately successful. He died in 1916 at the age

Jan 1, 1982

By Edmond Donovan, O&apos

With the increase in complexity of conveyor systems and the inexorable demand for more capacity and availability, the ability of designers to accurately predict performance is becoming increasingly important. Dynamic analysis techniques for modelling wave mechanics in conveyors are finding increased applications. Unfortunately, the reliability of these modelling techniques is dependent on a thorough understanding of the performance characteristics of the components that make up the system. These characteristics include such things as response rate, hysteresis and efficiency. Some parameters such as idler rolling resistance can be easily quantified by measurement. More complicated drive parameters are difficult or impossible to measure, particularly in large horse-power equipment. One method of quantifying or at least bracketing the performance envelope of complicated pieces of equipment is to construct a mathematical model. This can be done quickly and economically, and the influence of various parameters can be determined. This paper sets out to demonstrate by example the usefulness of this technique in gaining an understanding of complicated systems. The model is developed and the influence of a number of variables are examined.

Jan 1, 1996

The concept of optical basicity is now being widely applied to rationalize and predict the chemical properties of slags. It employs signals from a probe ion to indicate the "state" or relative "freeness" of the oxygen ions in the slag, and hence provides a fundamental measure of slag basicity. Values of the bulk optical basicity of a slag can either be measured spectroscopically in samples of transparent solid glassy slag, or calculated very simply and accurately from the chemical composition of the slag. The way in which these values vary in different systems are illustrated as isoptical basicity lines on ternary diagrams, and the patterns obtained are discussed. The difficulties associated with assigning optical basicities to the oxides of transition elements are also reviewed briefly. It has been demonstrated that these bulk values can be used as the basis of a flux control model for steelmaking to achieve end-point chemistries much more reliably. Furthermore, several correlations have been established which allow the prediction of the slag-metal distribution of various solute elements. This approach has been used to predict the phosphorus distribution at the end of steelmaking, and also to custom design fluxes for the ladle and tundish to aid in the control of residual elements in the steel such as suphur, hydrogen and nitrogen.

Jan 1, 1990

By F. N. Kissell, H. K. Sacks

This study examines the accuracy of area sampling for measuring the dust exposure of mining machine operators in coal mines. The specific objective of this research was to find locations where an area sampler might work better than earlier studies have indicated. The results show that fixed-location area sampling cannot accurately predict the dust exposure of a machine operator, even when the best fixed location is sought, the fixed location is quite close to the operator and the bias due to the dust concentration gradient is corrected. Industrial hygienists have known for many years that area sampling is unsuitable for measuring air contaminant exposures in the workplace. Near contaminant sources, the dilution air and the contaminants are not evenly mixed. Therefore, when workers are near contaminant sources, exposure measurements must be taken from the worker's breathing zone to be accurate.

Jan 1, 2003

By Gene E. McClelland

4.1 INTRODUCTION Metallurgical testing is required on any ore, waste or tailing material before amenability to heap leaching treatment can be accurately determined. The amount of testing necessary to determine and optimize heap leach amenability and conditions for any feed material is determined by the ore type, grade, size of deposit, and the overall financial commitment a particular company is willing to make. It is very easy to either "under-test" or "over-test" a particular deposit or feed material. A good heap leach testing program will provide the necessary design information without "bogging" down in too much detail. The three basic stages of ore testing include: • Preliminary--bottle roll tests and small column percolation leach tests; • Detailed--larger column tests, different feed sizes, and feed size optimization; and • Pilot-Scale--Column or field heap. The following subchapters describe the stages of heap leach testing necessary to put a viable property into production. 4.2 PRELIMINARY METALLURGICAL TESTING The two types of preliminary tests that can be undertaken include bottle roll or cyanide agitation tests and small column leach tests. Each of these is described below, along with information about what the test results provide in terms of furthering knowledge of ore qualities. 4.2.1 Bottle Roll Tests The first step in a heap leach testing program is preliminary bottle roll tests at a relatively coarse feed size to determine the ore's leachability. Bottle roll tests are usually conducted on reverse circulation drill cuttings samples (= 1/4 inch (6 mm) in size). Bottle roll tests provide initial information about precious metal recovery, recovery rate, and reagent requirements. The recovery results are to be considered maximums because attrition grinding creates fines and liberates recoverable values that would not be liberated from the material during a static leach. Bottle roll tests can be conducted on up to 25 lb (11 kg) ore charges of up to two inches (50 mm) in size. However, the coarser the feed the more attrition grinding occurs during rolling.

Jan 1, 1988

By Christian Jeger

After recalling the principles underlying fire-damp flow in coal and the essential part played by permeability (Darcy's law) in the intensity of flow, a few theoretical and experimental results showing the possible causes and range of variation of coal seam permeability are presented. The study of firedamp flow into gate roads and of residual gas concentrations in coals as a function of the gate road distance, shows that, in French mines, coal seams in a virgin zone have a very low permeability, but they become permeable when destressed by mine workings, in particular overlying workings. This is the case not only in the fractured zone round a mine excavation in which the gas is released, but also beyond the limit of the zone degassed by the excavation. In the floor of flat seams, upward boreholes can be used for predegassing seams that have been relaxed but not degassed, as the result of an overlying coal face. In this respect we shall describe the predrainage experiment carried out in the Marienau district of the Lorraine coalfield. Predrainage can contribute considerably to the reduction of methane emission in the ventilation of newly opened longwall faces. But to be efficient and useful predrainage should meet certain conditions.

Jan 1, 1980

By M. Leduc, M. Backens, M. E. Smith

An emerging technology for sustainable development in mineral development, tailings co-disposal™ is a means to dispose of conventional mill tailings within the void space of a waste rock dump. This method, while not applicable to all tailings systems, can offer a significant reduction in short-term and long-term environmental liability, land disturbance, closure costs, and negative public image. The concept is to dispose of the tailings simultaneously with the waste rock or leach ore, controlling the ratio of tailings to coarse material such that the resulting dump is stable both physically and chemically. The resulting co-disposal dump can be as safe as a conventional waste dump. In the case of alkaline tailings in a potentially acidic waste dump, the acid production can be buffered or slowed.

Jan 1, 2004

By G. Milojcic

The opencast lignite mines in the Rhinel and Di strict/Germany are equipped with the continuously operating bucket-wheel excavator - conveyor - spreader system. High productivity requires large capital expenditure on main mine equipment as a precondition for efficient mining. In operation, the technology employed leads to a particular cost pattern which is quite different from that of other industries. Labour, auxiliary equipment and mobile vehicles are the most important cost factors which can be controlled by the management. In this respect information processing computer systems are used for systematical and interpretable representation of operational tasks and jobs both in planning and work preparation and in recording the operational data describing the work result and the use of resources. Another aim of the project is to relieve the middle management of major part of its administrative routine work. Job-related work planning and data recording open up various possibilities of analyzing the operational procedures. This analysis shows, for example, what auxiliary machines are particularly suitable for specific tasks, and working groups can be organized so as to take account of the different skills of individual employees for the sake of an optimal work product. Computerized information processing provides the decision makers with topical information related to their respective responsibilities. The mining department of the Hambach opencast mine is the first user of the PERMAD system. Due to the favorable evaluation it has been decided to introduce PERMAD in three other mines in the course of 1992/93.

Jan 1, 1992

By Bernard J. Scheiner, R. J. Wesely, Cy E. Jordan, A. D. Zunkel, Gary M. Wilemon

The U.S. Bureau of Mines has investigated leaching of off-grade western United States phosphate deposits using sulfuric acid (H2SO4) in the presence of methanol. A combination of ore grinding followed by sulfuric acid in methanol leaching extracted over 80 pct of the phosphate. This leaching technique significantly limited the dissolution of iron, aluminum, and magnesium impurities and produced a phosphoric acid that was well within the desired limits for these impurities. The dissolution rate of phosphate rock in the sulfuric acid methanol system was significantly slower than the conventional aqueous sulfuric acid dissolution. Initially dissolution of the phosphate particle was controlled by the amount of surface area available to the sulfuric acid reactant. After about an hour of leaching, phosphate dissolution was slowed almost to a stop, limited by the rate of ion diffusion through the calcium sulfate barrier layer surrounding the unreacted phosphate particle. Consequently, the ore was ground below 38-pm to increase the particle surface area available and improve the overall dissolution kinetics and extraction. In addition, the sulfuric acid was slowly metered into the reactor to prevent any excess heat from the mixing of acid with the methanol from simulating the formation of methyl sulfates. The formation of methyl sulfate ultimately lowered the overall phosphate extraction by exhausting the supply of available sulfate ions for the dissolution reaction. Subsequent evaporation of the methanol from the product acid decomposed the methyl sulfates leaving undesirable sulfate ions in the product acid. Slow acid addition to the reactor provided high phosphate extraction and limited the amount of sulfate in the final product acid.

Jan 1, 1993

By W. E. Hale

The Institute of Applied Geology (I.A.G.), a joint project sponsored by the University of the Philippines, the Philippine Bureau of Mines and the United Nations Development Programme (Special Fund) undertook during 1965-66 a number of field projects in Northern Luzon Island (Fig. 1). These field operations were of a two-fold nature: during approximately six weeks an orientation survey, involving, geology, geophysical surveying, and geochemical prospecting, was conducted over a partly developed, disseminated copper deposit at Sto. Nino (Fig. 1). The Sto. Nino study demonstrated (Hale and Govett, 19677 Govett and Hale, 1967) the feasibility of geochemical prospecting relative to other exploration methods for the location and delineation of disseminated copper occurrences, the major source of metal in the expanding mineral industry of the Philippines. A vein occurrence at Suyoc (Fig. 1) was similarly subjected to

Jan 1, 1968

By Norman Weiss

The design of an ore-processing plant requires a high degree of cooperation among the geological, mining and metallurgical engineers. The purposes of this section are to provide mill design information most commonly needed by the mine planners and management, and to summarize the kinds of information that must be supplied to the mill designer to enable him to provide a suitable plant. 28.1-BASIS OF DESIGN NORMAN WEISS 28.1.1-SIZE OF PROJECT The size of the project generally is expressed in tons of ore milled per day, but company policies differ widely in this respect, a few adhering to this set figure but many expecting the design capacity to be liberal. To some companies that follow the latter practice a 25,000-tpd milling plant that cannot reach 30,000 or 35,000 tpd without major additions is a disappointment. The engineers responsible for planning the operation must know what is intended. Equally important are operating schedules, particularly the number of operating days per week and shifts per day for ore delivery, crushing, grinding and processing. Hours per shift also are an important factor in design, particularly in the case of ore delivery, and the percentage lost time expected for maintenance, cleanup, inspection, preventive maintenance shutdowns, and many others, has to be taken into account. Frequently, the ore deposit will vary in hardness and grade to such a degree that the capacity of the mill will vary widely from year to year and cannot be expressed as a fixed figure. but rather as an average over a calculable period, such as 10 yr, or as a specific daily capacity over the first 2 yr, next 4 yr, etc., or some similar projection. A good example of this kind of situation is Asarco's Mission operation 15 mi southwest of Tucson, Ariz., described by Weiss and Vincent.' where the major rock types differed widely in hardness, so that a change in the proportion of these types in the mill feed could have imposed difficult design problems if a constant tonnage rate had been expected. 28.1.2-EXPECTED LIFE The life of a mining operation depends upon the size of the deposit and the rate of mining. The latter is a policy decision based on an economic analysis which, in turn, is based on many factors.' From the point of view of mill planning, the estimated life of the operation determines many criteria and affects many decisions on strength of structures and degree of protection of men and materials, the quality of the materials-handling and process machinery, the type of delivery systems for ore to the mill and concentrates to the smelter or market, and many others. Costly installations that may be excellent investments for a 20-yr project may be wasteful for a 5-yr operation. An example is the comparison of different methods of crushing or grinding. Take single- vs. multiple-stage grinding-the former may be found more economical for a 5-yr operation, and the other, more costly, method more economical for a 15 to 20-vi- life.

Jan 1, 1973

By T. M. Karppanen

Shotcrete is used both in mining and tunnelling for supporting rock. Both dry and wet processes are used for shotcreting, and each produces good quality concrete. Thick shotcrete layers function as a concrete arch supporting loose rock. Thin layers prevent relative movements between the blocks of rock due to the fine concrete in the cracks and joints. No progressive fallouts occur since key blocks are kept in place. Rockbolts are often used together with shotcrete in order to create an arch around the tunnel making the rock to support itself. Shotcrete is sprayed either manually or by means of mechanized equipment. The aim is to spray required the layer thickness in one pass and with a minimum amount of rebound. Sprayed layers should also be as even as possible. The tip of the nozzle should be moved in circular movements. The shotcrete should be sprayed at a right angle to the rock surface; the optimum distance from the nozzle to the application surface must be found by experiment. In a mine shotcreting forms a materials handling system which includes batching, transporting and the actual shotcreting. The equipment used vary depending on the spraying method, mining conditions and the location of the mine. Shotcreting is performed with a purpose-designed unit in most cases where operation is mechanized. The shotcreting unit consists of a spraying machine with a hose and a nozzle, an additive system, a manipulator and a mobile carrier with all the necessary controls and accessories. The key element of the unit is the shotcreting manipulator which holds and guides the nozzle as required. The manipulators currently in use are hydraulically operated, and their remote controlled booms are also often called robots. These manipulators have enabled users to increase their output capacities, to remove the operator away from unsafe areas and from the influence of dust and rebound. Furthermore, they improve the shotcrete quality and reduce shotcreting costs.

Jan 1, 1989

By John G. Whellock

Over the last decade new constraints have emerged in the metallurgical industry which challenge the very basis of design and construction of smelters and refineries. The hackneyed designs of the past will not now serve the revised objectives of: - minimal cost of ownership - improved productivity - reduced energy consumption - greater flexibility in feed or product - minimized recycle loops and by-products - more stringent environmental control The impact of these constraints has been channeled constructively into process intensification - exemplified at the large scale by the introduction of flash smelting technology and at the smaller scale by the application of rotary process vessels and converters for augmented heat and mass transfer. The design of successful process plants depends on the understanding of the unit operations and their interactions which is the discipline of systems engineering. With the advent of increased process intensification the importance of a systems approach to process design and implementation becomes paramount. This paper describes the systems approach with specific reference to medium scale projects in the primary and secondary metallurgical industry.

Jan 1, 1985

By Thomas E. McNider, Robert C. Howell, J. W. Stevenson

As of the end of October, 1990, there have been eight spontaneous combustion heatings reported to the Mine Safety and Health Administration (MSHA) at the No. 5 Mine. A reportable heating is defined as one with elevated surface temperatures of the strata greater than ll°C (20°F) above the ambient. The first heatings occurred in November, 1986. The last three heatings were located near the start line of "K" Longwall Panel in the southwest where mining had begun on May 21, 1990. All of the heated areas had been at the time of detection, or were sometime areas of high ventilation pressure differentials or high air velocities. All of the heated areas are in an expanded seam with associated, extensive floor heave which is characteristic of all of the expanded seam thus far exposed by development mining in the No. 5 Mine. Jim Walter Resources, Inc. submitted a plan to MSHA which would reduce the potential for spontaneous combustion at the No. 5 Mine by reducing high air velocities and high pressure differentials in the gob. These two parameters are the dynamics which feed oxygen to the pyrite, thus initiating the oxidation process which can lead to spontaneous combustion. No. 5 Nine operated under this plan for 29 shifts before mining on the longwall was stopped by MSHA and the Spontaneous Combustion Control Plan revoked. This paper briefly details the geology of the expanded seam and history of the heatings. It also gives a retrospective view of the operation under the Spontaneous Combustion Control Plan and reasons why it was revoked by MSHA. Currently, knowledge of the reaction process of the footwall coal seams at No. 5 Mine and other potential spontaneous combustion seams in the United States is limited. There are techniques used in the United States and other countries to minimize the potential for heatings in spontaneous combustion prone areas. The No. 5 Mine has attempted to utilize this current technology. However, in deep, gassy mines the problem is complex.

Jan 1, 1991

By Lester E. Sutton

The past three years have been the most difficult period in the modem history of the barge line industry. Having withstood the fuel cost run ups of the late 1970s and the imposition of inland waterways system user fees, the industry has struggled more recently with an overcapacity problem of unprecedented proportions. In responding to these challenges, the industry is undergoing important long-term structural changes, with the various individual barge lines pursuing an array of divergent long-term business strategies. This paper discusses these strategies in terms of their probable impact on the industry as a whole, and argues for the survival of a competitive number of large, national, independent barge lines as the outcome most in keeping with the long-term interests of the mining and minerals industry.

Jan 1, 1986

By Ian S. Ross

The emphasis in this paper is focussed on dry bulk car go or the same cargo converted to slurry form for ease of handling. It is this type of cargo which is of primary interest to the mining industry whose products it represents to a large extent. The design problems associated with the handling of liquid bulk and of unitized cargo are sufficiently different to be treated separately. This paper then will consider the port handling facilities of dry bulk commodities as part of a system consisting of (1) the movement of products to the port area, (2) the movement of products out of the port area and (3) the handling of products within the port area itself. Expanding within these three components of the total transportation system, the paper will discuss the various difficulties encountered concerning the mode of transport into the port area and the method of handling product from transport to stockpile. Secondly, and before discussing the port area itself and in order to complete the conditions under which the port must operate, the movement of products out of the port area will be covered dealing with such items as the accommodation of the ship, problem associated with the very large ships, optimum berthing arrangements and the need to consider the various costs associated with the vessel. Finally the paper will review within the port or interface area such items as stockpiles, method of recovery, inventory control and an outline of the common design difficulties encountered. In all cases the relative influence on the total transportation costs of each of these dry bulk transportation system components will be considered.

Jan 1, 1972

By R. J. Hamilton

Considerable success has been achieved in recent years in reducing concentrations of respirable dust in UK coalmines; thankfully this has been reflected in significant improvements in pneumoconiosis statistics. However, continuous increases in mechanisation and in rates of working, accompanied by more stringent dust regulations, necessitate the introduction of improved and often more sophisticated methods of dust control. The increased application of exhaust ventilation and air cleaning techniques seems likely to be the most important step forward in dust control in coalmining during the next few years. A typical drivage in a UK mine is not more than 16 ft (5 m) wide and is supported by steel arches. In the past the majority of such roadways were driven by the conventional means of blasting and loading. Since the main purpose of ventilation was to clear blasting fumes and to provide adequate scouring of firedamp, not surprisingly it was usual to have 'forcing' ventilation. Present systems of operation are making increasing use of heading machines and it is not surprising that well over half of all drivages are now ventilated by exhausting methods. (1) The great majority of coal getting in the UK is from longwall faces, usually in seams of less than 6 ft (1.8 m) thickness. Although this method of min¬ing has many advantages in terms of safety and productivity, dust control is made more difficult because of the need for men and machines to work in the same ventilation stream. Much emphasis has been placed on improving the design and methods of opera¬tion of longwall machines to reduce dust production, and improved ventilation has provided greater dilution, and in consequence, reductions in dust concentration. On all machines water is now fed through internal galleries to the cutting elements. Unfortunately, one of the difficulties associated with mining thin seams is that there is some unavoidable cutting in the roof or floor and in many seams there are faults to be traversed and dirt bands which must be removed with the coal. Under these conditions it appears that the use of exhaust ventilation and air cleaning offers the most promising way of achieving satisfactory dust conditions.

Jan 1, 1980

By Nai-piu Tong, Huaren Dou, Sui-shun Poon, Leslie H. Swann

This paper mainly presents geotechnical aspects of a vehicular underpass that was built across under one of the busiest highway in Hong Kong. The underpass, in a section comprising an upper arch and invert slab, is approximately 6.2 m high at crown and 15 m wide at the base level with 40 m in length. The construction inevitably requires the underpass to be built using tunneling techniques to keep the overheadroad traffic uninterrupted. The paper will describe the design of temporary support for the tunnel excavation that was provided by an arch formed with horizontal pipes, which were installed using the pipe ramming method. The pipes arch was then supported by steel ribs founded on rock or mass concrete. Permanent lining design consideration was also discussed. Results of numerical modeling of the soil-structure interaction using the 2-D computer code FLAC will be presented. The predictions have been compared to the monitored results during construction, which indicates the prediction is in reasonable agreement with the measurement in the light of ground movement. The construction challenges will also be discussed in the paper.

Jan 1, 2003

By J. Mark Richardson, Mary S. Sredanovic

The use of computerized steady-state process design balance systems has reduced engineering man-hours by as much as 80 to 1 over conventional manual balance calculations. The popularity and widespread use of such systems by the minerals industry increased dramatically in recent years, but these systems were mainly available for use on large main frame computers and lacked graphic capability. A considerable amount of engineering manpower was required to "conceptualize" a process flowsheet. After sketching the flowsheet layout, the engineer was still faced in most cases with a time-consuming reduction of his graphic concept of the problem to numerical data required by the balance program. Furthermore, output from these early systems was usually in tabular numeric form. The computer results then were transferred manually to a pictorial representation of the process. A new process design balance system has been developed which runs on a popular and inexpensive personal computer. This system combines modern Computer Aided Design (CAD) principles and the graphics capabilities of the personal computer to allow the process engineer to define a flowsheet problem while creating a process flow diagram. The system provides balance results in both pictorial and tabular form. Past, present, and potential applications of the system are discussed, and an analysis of the cost benefits derived from its use are provided.

Jan 1, 1985

By Richard L. Bullock, Bruce Kennedy

Vince Lombardi once said, "Practice doesn't make perfect, perfect practice makes perfect." When it comes to building a mine that will operate at the optimum level for the set of geologic conditions from which it was developed, Lombardi's remark might be paraphrased to describe the problem: planning won't guarantee the best possible mine operation unless it is the best possible mine planning. Any sacrifice in the best possible mine planning introduces the risk that the end results may not reach the optimum mine operation desired. This section addresses many of the factors to be considered in the initial phase of mine planning. These factors have the determining influence on the mining method, the size of the operation, the size of the mine openings, the mine productivity, the mine cost, and, eventually, the economic parameters used to determine whether or not the mineral reserve even should be developed. A little-known fact, even within the metal-mining community, is that room-and-pillar mining accounts for most of the underground mining in the united States. According to a 1973 study on noncoal mining (Anon., 1974), more than 76% of the producing mines [of over 1089 t/d (1200 stpd) capacity] produced approximately 70 000 000 t (77,000,000 st) or 60% of the nation's underground tonnage of material by room-and-pillar mining. That same year, 96.8% of the nation's under- ground coal mines produced 262 950 000 t (289,911,000 st) of coal extracted from room-and-pillar mines (Anon., 1976). Thus, nearly 333 000 000 t (367,000,000 st) of the United States' raw material is produced from mines using some form of the room-and-pillar mining system. Because approximately 90% of all mining in the United States is done by some variation of room-and- pillar mining, it is appropriate to give special emphasis to the effects of the various elements of mine planning on room-and-pillar mining. The relationship of these elements to other mining methods will become apparent as the elements are described in later sections herein. TECHNICAL INFORMATION NEEDED FOR PRELIMINARY MINE PLANNING Assuming that the reserve to be mined has been delineated with diamond-drill holes, the items listed in the following paragraphs need to be established with respect to mine planning for the mineralized material. Geologic and Mineralogic Information The geologic and mineralogic information needed includes the following: 1) The size (length, width, and thickness) of the areas to be mined within the overall area to be considered, including multiple areas, zones, or seams. 2) The dip or plunge of each mineralized zone, area, or seam, noting the maximum depth to be mined. 3) The continuity or discontinuity within each of the mineralized zones. 4) Any swelling or narrowing of each mineralized zone. 5) The sharpness between the grades of mineralized zones within the material considered economically minable. 6) The sharpness between the ore and waste cutoff, including whether this cutoff can be determined by observation or must be determined by assay or some special tool; whether this cutoff also serves as a natural parting resulting in little or no dilution, or whether the break between ore and waste must be induced entirely by the mining method; and whether or not the mineralized zone beyond (above or below) the existing cutoff represents submarginal economic value that may be- come economical at a later time. *7) The distribution of various valuable minerals making up each of the minable areas. 8) The distribution of the various deleterious minerals that may be harmful in processing the valuable mineral. 9) Whether or not the identified valuable minerals are interlocked with other fine-grained mineral or waste material. 10) The presence of alteration zones in both the mineralized and the waste zones. Structural Information (Physical and Chemical) The needed structural information includes the following: * 1 ) The depth of cover. 2) A detailed description of the cover including: the type of cover; * the structural features in relation to the mineralized zone; * the structural features in relation to the proposed mine development; and * the presence of and information about water, gas, or oil that may be encountered. 3) The structure of the host rock (back, floor, hanging wall, footwall, etc.), including: * the type of rock; * the approximate strength or range of strengths; * any noted weakening structures; * any noted zones of inherent high stress; noted zones of alteration; the porosity and permeability; * the presence of any swelling- clay or shale interbedding; the rock quality designation (RQD) throughout the various zones in and around all of the mineralized area to be mined out; the temperature of the zones proposed for mining; and the acid generating nature of the host rock. 4) The structure of the mineralized material, including all of the factors in item 3 plus: * the tendency of the mineral to change character after being broken, i.e., oxidizing, degenerating to all fines, recompacting into a solid mass, becoming fluid, etc.; * the siliceous content of the ore; the fibrous content of the ore; and the acid generating nature of the ore. Economic Information The needed economic information includes: *1) The tons of the mineral reserve at various

Jan 1, 1982