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By T. J. Jr. DeMull, F. G. Miller, J. P. Matoney

For some time a need had existed in the minerals processing field for a relatively efficient separator that would treat high tonnages of particles in the intermediate size range, i.e., those particles too large for froth flotation and too small for conventional gravity-type separa¬tors. Among those devices developed to meet this need are the centrifu¬gal specific gravity separators. These devices employ centrifugal acceleration to assist gravitational acceleration in separating light¬density minerals from heavy-density minerals. In the category of centrifugal specific gravity separators are the heavy-media centrifugal separator and the water-only cyclone. The two major centrifugal heavy-media separators, i.e., the heavy-media cyclone and the DynaWhirlpool, as well as the water-only cyclone, are discussed in terms of: design features, operating variables, operat¬ing data, and flowsheet design criteria. Examples of plant applications are given in the field of coal processing as well as the processing of other minerals such as iron ores, potash, and tin. Finally, the subject of the staging of centrifugal separators and their use in combination with other separators is discussed. PRINCIPLES For coarse sizes of minerals, efficient specific gravity separations have been possible for many years with open-bath vessels using the natural settling velocity or buoyancy of the particles. These bath ves¬sels process ore by utilizing micron-size solid particles suspended in the slurry fed to the separator. The inclusion of these particles in the slurry increases the effective density of the separating fluid to allow particle separations to be made at densities greater than that of water. However, if vessel size is to remain within economical limits, the particles processed in the bath vessel must have high settling rates in a IG gravitational field. Because of this requirement, heavy¬medium bath vessels are usually restricted to processing +V4-in. sizes. To extend efficient specific gravity separation to smaller sizes, the gravitational acceleration of particles is replaced by centrifugal acceleration. The settling of a small particle in a fluid in a centrifugal force field is similar to that found in a static bath except that the acceleration due to gravity, g, is replaced by a centrifugal acceleration where v, is the tangential velocity at radius r: V=kdm(P-P,), V'. (J) µ In more practical terms where the particles settle in a suspension of finer particles comprising the heavy media and with an effective suspension density p" V = kdm' (P P ) . v'. (2) U r To date the most effective use of this principle has been obtained with devices that rotate a liquid or suspension within a stationary enclosure in order to create centrifugal force. Cyclones are the most common devices used for this purpose, because they generate centrifu¬gal forces far greater than the force of gravity and therefore not only have high capacities but can treat finer sizes than bath-type vessels can. The two main types of cyclones used by industry are the heavy-media cyclone and the water-only cyclone. Also quite widely used is the DynaWhirlpool, which, though based on the same princi¬ple, differs in design from the conventional cyclone. HEAVY-MEDIA CENTRIFUGAL SEPARATORS Like the bath vessels, the heavy-media centrifugal separators em¬ploy media composed of micron-size particles suspended in water. However, the centrifugal force generated in these separators accentu¬ates the difference in settling rate between particles of different density and thus makes possible separations of finer size particles than can be treated in bath vessels. The two most common heavy-media centri¬fugal separators are the heavy-media cyclone and the DynaWhirlpool. Heavy-Media Cyclone Although cyclones were originally developed for use as classifiers or thickeners, it was later found that they could also effectively serve as heavy-media separators.63. 64, Design Features Fig. I I is a schematic of a typical cyclone developed to serve for any one of the following purposes: as a classifier, thickener, or specific gravity separator. The cyclone consists of a cylindrical section joined to a conical section, usually having an included angle of between 14° and 25°. Feed enters the cyclone tangentially through an orifice attached to the cylindrical section. The overflow orifice is located in the base plate of the cylindrical section. The vortex finder, a tube attached to the overflow orifice, extends into the cyclone from the base plate of the cylindrical section. The underflow orifice is located at the apex of the conical section. As some medium together with mineral particles is fed through the feed orifice, a vortex with a hollow air core extending from the overflow to the underflow orifice forms in the cyclone while hollow spray discharges form at each of these orifices. Under the influence of the centrifugal force, high specific-gravity particles move through the medium to the wall of the cyclone and descend in a spiral flow pattern to the underflow orifice. Those particles in the feed stream, lower in specific gravity than the feed medium, follow the major portion of the flow to the center of the core where they are caught in the high-velocity upward central current and are carried out through the overflow orifice. Fig. 12 shows a family of curves that illustrates how materials of varying specific gravity are recovered by a cyclone. Since the specific gravity of the medium is 1.40, the particles of 1.40 sp gr actually act as part of the medium and, regardless of the particle size, split between the underflow and overflow of the cyclone in proportion to the volume split of the medium. Particles higher in specific gravity than 1.40 are recovered in the underflow of the cyclone at increasing rates as the difference in specific gravity increases and the particle size increases. Particles lower than 1.40 in specific gravity are dis¬charged through the overflow orifice at increasing rates as the specific¬gravity difference increases and the particle size increases. However, all the curves originate at the fluid-flow ratio point for the finest particles of any gravity. The fluid-flow ratio is defined as the ratio of the rate of fluid flowing from the underflow to the rate of fluid

Jan 1, 1985

By Lok W. Home

HISTORICAL DEVELOPMENT At the beginning of the 1970 more than 90% of pro¬duction drilling in open-pit mining was obtained with the use of rotary drills. At the same time, the use of rotary drills underground was virtually nonexistent. Several progressive mines began questioning whether efficient rotary drilling could not be adapted to under¬ground applications. In the late 60s and early 70s at¬tempts to use rotary drilling were made by Inco, Ltd. and Meramac Mining, utilizing drill units designed as a combination of small diameter raise drills and rotary drills. These initial attempts were unsuccessful primarily due to the immobility of the equipment and low degree of dependability. At the same time, experiments con¬ducted by Inco and Meramac using large diameter in¬the-hole percussion drills produced immediate encourag¬ing results. The first serious attempt at utilizing rotary drilling underground was made by Texasgulf in 1971 at Kidd Creek mine in Canada. An experimental stope 15 m wide, 36 m long, and 30 m deep was drilled using 200-mm diam rotary drilled holes. The machine used was a Robbins' Model 11 D. The Model 11D is a rotary drill designed to drill drain and utility holes under¬ground. The results of this test revealed that large diameter blasthole stoping was feasible. The indications were that the major areas of concern-fragmentation and pillar damage-could be controlled and a new low-cost stoping method resulted. At this point Texasgulf and Robbins entered into a cost-sharing agreement to develop a rotary drilling unit specifically for this stoping method. Drilling of the trial stope required pouring a concrete pad at each drill hole plus anchoring the derrick to the pad by means of rock¬bolts. Utilization expressed in terms of drilling time to available time was less than 20%. It was obvious that the required unit had to be mobile, setup time had to be minimized, and rod handling had to be improved. The first attempt at an underground production-type rotary drill consisted of mounting the basic 11D on an air-powered Eimco 600 Series crawler. The crawler was equipped with four screw-type leveling jacks operated by an impact wrench; rod rack facilities for 60 m of 1.2-m long, 170-mm diam rods; a rod handling facility consisting of an air-lift swing boom and a pipe loader which automatically positioned the rod under the float¬ing drive box; a turnbuckle anchor system requiring in¬stallation of four rockbolts in the drill drift floor and placing turnbuckles from the rockbolts to the corners of the drill; and trailing power units containing the electrics and hydraulics required to operate the drilling unit. Operation of this unit revealed a vast improvement in drilling time but some problem areas remained. The anchoring system required secondary drilling and proved difficult to adapt to the localized irregular floor surface and the trailing power unit was difficult to maneuver and was causing unnecessary maintenance costs. Addi¬tional improvements resulted in the standard 11MD rotary blasthole drill. DESCRIPTION OF UNIT This standard model consists of the drill and the power unit integrally mounted on an air-powered crawler chassis (Fig. 1). The rod-handling and leveling¬jack systems are similar to those used on the prototype Texasgulf unit and remained relatively the same. The anchoring system has been replaced by the roof jack, initially screw type but now hydraulically actuated. The anchoring system is necessary as the drill exerts more thrust than the weight of the unit. During operation it is found that the two front jacks adequately anchor the machine. The complete unit's operational dimensions are: height, 3.2 m; width, 2.1 m; and length, 4.2 m requiring a 3.5- x 3-m drill drift. The derrick is pivot mounted and raised and lowered by a hydraulic cylinder reducing the transport height to 2.5 m. The derrick is also turnbuckle mounted to allow dip angle drilling from vertical to 0.52 rad (30°) off vertical. The drill derrick can be mounted to allow dip axis to be parallel or perpendicu¬lar to the long axis of the unit. The standard rods are 1.2 m in length and are contained in racks of 10 which can be removed for transportation. Drilling torque is produced by a 56-kW (75-hp) electric motor driving a double vane pump which in turn drives a Robbins high torque gear-type hydraulic mo¬tor. In addition to the hydraulic motor, the machine drive system has a 3.9:1 planetary gear reducer. This results in six speeds with the following minimum and maximum range: Maximum rpm 112 Torque 3725.5 J (380 kg m) Minimum rpm 10 Torque 14 706.0 J (1500 kg m) Drilling thrust is provided by two bootstrap cylin¬ders pressurized by a variable volume pump. This sys¬tem can produce from 0 to 267 231.2 N thrust in either direction. Bailing air (approximately 17 m3/min at 382.2 to 490.3 kPa) is introduced through a rotary swivel pass¬ing directly through the center of the drive system to keep efficiency and losses to a minimum. Hole annulus is also kept to a minimum by using 170-mm rod when drilling a 200-mm hole. A small amount of water is added to the air for dust suppression. The machine has to be supplied with air, water, and electrical power. It does not appear feasible to provide an integral air power supply with the rotary drill such as is done with open-pit rotary drills. Air is required and installed during the primary development of the drill drift. The machine can be adapted to drill hole diameters from 170 to 228 mm for blasthole operations. The drill string consists of the rotary bit, crossover stabilizer, roller stabilizers, and standard rods. The operator controls speed and thrust pressure from an operating console which can be mounted in a conve¬nient position. Mount Isa has made an adaption to the operating console to permit remote collaring of drill holes in an uncleared floor.

Jan 1, 1982

By Noel A. Genco, Edward F. McCarthy, Robert J. Piniazkiewicz

Talc, when it can be isolated as a pure mineral, has a composition of 63.36% SiO2, 31.89% MgO, and 4.75% H,O. However, as a commercial commodity less than half of all talc sold has a purity exceeding 95%. Nevertheless, impure talc products find a multitude of uses and few substitutes in many industrial applications. Minerals commonly associated with and sold together in talcose mixtures are chlorite, dolomite, mica, magnesite, and tremolite. Steatite was originally a mineralogical term applied to pure talc. Today steatite usually denotes a high-talc ceramic body used as an electrical insulator, and the talc used in such applications is known as steatitic talc. Steatitic talc generally is found in massive, high- purity form and in the past was machined directly into electrical parts, which rarely is done today. Impure varieties of massive or block talc are still commonly termed soapstone. Soft massive talc, suitable for crayon manufacture, has been referred to as French chalk. It has been common practice to discuss talc, soapstone, and pyrophyllite under the same general heading. In the case of soap- stone and talc this is natural, since many different types of platy, soft minerals exhibiting a high degree of lubricity (commonly referred to as slip) have been grouped together and called soapstone or talc. Further, there are mineralogical relationships between pyrophyllite and montmorillonite and between talc and hectorite. Thus, talc and pyrophyllite are sometimes referred to as clay minerals. When finely divided talc or pyrophyllite are combined with water in proportions to make a slurry, the end product does have the appearance of a clay-water mixture. Soapstone for utensils and ornaments was mined by prehistoric Native Americans on Santa Catalina Island, CA. In the mid- 1800s soapstone from deposits along the western foothills of the Sierra Nevada Mountains was used by white settlers as building and ornamental stone and for the linings and foundations of furnaces (Anon., 1956). Previous to 1916 the annual recorded production in California did not exceed 1.8 kt. From 1912 to 1918, however, output rose sharply when the Talc City, Western, and Silver Lake mines were put into operation. From 1916 to 1935, the state's annual output of talc was in the range of 8.2 to 18.2 kt. There was another increase in the mid-1930s, when the use of talc in wall tile grew rapidly. In California the production of talc grew to 57.2 kt in 1943. The post-war building boom helped California's production to grow to 108.9 kt in 1951. Production stood at 149.8 kt in 1968, 13 1.6 kt in 1969, 167.9 kt in 1970, and 140.7 kt in 1972. In the mid-seventies and early eighties, however, California's talc production declined to approximately 20 kt in 1992 due to reserves exhaustion and the material's tremolite content, which raised significant product liability concerns at the time. Talc mining in New York State dates from about 1878, when a Colonel Palmer and associates opened the first commercial talc mine on the Nelson Freeman farm near Talcville. In 1893 this operation was sold to International Pulp Co., which changed its name to International Talc Co. in 1944. The company was acquired by R.T. Vanderbilt Co., Inc., in 1974. A new talc operation, Gouverneur Talc Co., owned by R.T. Vanderbilt Co., began operations near Balmat, NY, in 1948. The initial capacity of the processing plant was 180 tpd (Gillingham, 1950). Subsequent expansions have increased capacity to more than 725 tpd. Talc was discovered in the area of Johnson, VT, in 1902, marking the birth of the Vermont talc industry. American Minerals Co. initiated plant operations at Johnson in 1904. MagnesiaTalc Co. opened a plant at Waterbury, VT, in 1913 (Burmeister, 1963, Trauffer, 1964) and acquired American Minerals Co. in 1923. The Waterbury mine had a long history as a producer of talc crayons. Eastern Magnesia Talc Co. was formed in 1924 by the merger of Eastern Talc Co. and Magnesia Talc Co. In 1956 Vermont Mineral Co., with a talc mine at Hammondsville, VT, was also acquired. A new plant was then built at Gassetts to produce roofing products. In order to produce high grade products for several markets, Eastern Magnesia Talc Co. began operating a modem froth flotation plant at West Windsor, VT, in 1964. The ore was provided by the Hammondsville mine. This operation, known as Windsor Minerals, Inc., continued from 1972 to 1989 as a subsidiary of Johnson and Johnson Co. In 1989 Cyprus Minerals Co. acquired Vermont Talc, Inc. and Windsor Minerals, Inc.. thus consolidating the talc mining and processing operations at Hamm, Windham, Johnson, Chester, Ludlow, West Windsor, Troy, and Reading, VT, under one ownership. As the result of Cyprus' sale of its worldwide talc assets to the RTZ Group in 1992, its US talc operations are now part of Luzenac America, Inc.

Jan 1, 1994

By A. J. Richardson

Before the development of the underground stoping and mining costs can be considered, certain facts about the ore body, the proposed mine, markets, etc., must be known or determined. In the case to be studied, the zinc-lead mineralization occurred with a narrow vertically dipping structure of undetermined length and vertical extent. Exploration completed to date has revealed 6.5 mil¬lion st t of proven reserves. A further 820,000 st of in¬dicated reserves has been outlined and this tonnage is considered capable of being expanded by a factor of approximately four after more detailed drilling. After studying the market conditions and completing a very preliminary feasibility study, it was decided that production would be 730,000 stpy (or 2000 stpd) of ore. First year production would be at the rate of 1500 stpd. The main design criteria for the selection of the min¬ing methods are minimizing surface subsidence, maxi¬mum recovery of the ore body, maximum degree of grade control, maximum productivity, and safe working conditions. Two basic extraction systems are considered capable of meeting these requirements: mechanized cut¬and-fill stoping and sublevel long-hole stoping with filling. The primary development system of the mine has been designed to give maximum flexibility in stoping systems and layout and to permit changes if considered necessary as a consequence of actual production ex¬perience. At the present time, access to the mine is by a circu¬lar concrete lined vertical shaft, 16 ft diam, sunk to a depth of 1380 ft. Two exploration levels have been driven within the ore zone at depths of 165 and 1246 ft below the surface outcrop. The development to date had the objective of sampling the mineralization and produc¬ing detailed information on the outline of the ore body and the distribution and controls of zinc and lead values. In an attempt to satisfy the basic design criteria for the mine, it was decided that production would be best achieved by a combination of 40% sublevel long-hole stoping and 60% cut-and-fill mining. Costs of exploration and capital development of per¬manent underground facilities are normally written off over the life of a mine. Production expenditures, on the other hand, are of a temporary nature and are normally charged as and when incurred as an operating expense. Reasonably accurate predictions of mine production costs can be built up from engineering design and estimates of individual mine activities for ultimate inclusion in the comprehen¬sive data required for financial decision making. The simulated operations can be costed on a detailed basis in the form of a monthly operating budget. The budget format can be generalized or detailed, depending upon the scope of the project. However, ex¬perience suggests that a fairly detailed format has the advantage of assuring that all significant cost items are included. For underground costing it is suggested that the budget structure include five major cost centers (i.e. development, diamond drilling, ore extraction, hoisting/ transportation, and general mine expense). These in turn are detailed under numerous subheadings. The mechanism for compiling an operating budget will be illustrated. Because of its relative simplicity, ore extraction under sublevel long-hole stoping has been chosen for illustration. All other activities, simple or complex, can be estimated in similar fashion. BLOCK AND STOPE DEVELOPMENT Long-hole blocks, used where advantageous, will be up to 250 ft in height, depending upon the vertical con¬tinuity of the mineralization, and approximately 300 ft long. Drawpoints will be at 36-ft intervals and serviced by loading crosscuts driven from a footwall drift parallel to and close to the ore zone. Pillars between the stopes will be 50 ft wide. Stopes will be drilled off with vertical rings of blastholes drilled from sublevels approximately 60 ft apart vertically. This drilling will be done by percussion drilling machines (31/2 in.) mounted on a trackless drilling rig. Load¬haul-dump (LHD) equipment will be used to move broken ore from the drawpoints to the orepass connecting to rail haulage systems. On completion, long-hole stopes will be backfilled to prevent caving and to facili¬tate later pillar removal. From a planned stope layout, a forecast of produc¬tion and development is made in Table 1. Table 1. Block Tonnage and Stope Development Quantity Ore Waste Total ore block 375,000 st 2 stopes 310,000 st 1 pillar 65,000 st Access crosscuts, 4 at 100 ft 400 ft Drill sublevel drifts, 6 at 300 ft 1800 ft Stope raises, 3 at 250 ft 750 ft Undercut sublevel drifts, 2 at 300 ft 600 ft Loadout crosscuts at 35-ft intervals 550 ft 100 ft 3300 ft 500 ft Total development footage 3800 ft Tons per ft of development 987 st

Jan 1, 1982

By Roy M. Fleming, Christine B. New

The federal standard for limiting exposures to miners from radon daughters was reduced from 12 working level months (WLM) per year to 4 WLM per year in 1971. However, even at that time some researchers were concerned that the new limit would eventually be shown to result in an excess risk for lung cancer mortality in miners. The National Institute for Occupational Safety and Health (NIOSH) is now engaged in a comprehensive review of this topic. As the lead agency in this project, NIOSH has developed a work plan and established a work group to implement this plan. The procedure and specific considerations are outlined in the work plan for developing a comprehensive ionizing radiation standard recommendation for all miners, underground and surface. Such a recommendation will not only include estimates of health risks at various levels of exposure, but also appropriate recommendations for medical monitoring, sampling and analytical methods, sampling strategies, posting, engineering controls, personal protective equipment and recordkeeping. The work group has twenty-four members. Eleven members are NIOSH personnel representing six divisions of the Institute. The remaining thirteen members represent other federal agencies, specifically the Bureau of Mines of the Department of the Interior, the Mine Safety and Health Administration and the Occupational Safety and Health Administration of the Department of Labor, the Office of Radiation Protection of the Environmental Protection Agency, and the Bureau of Radiological Health of the Food and Drug Administration, Department of Health and Human Services. Several factors contributed to the decision to initiate this project. First, the efficacy of the current standard was considered. An initial study group was formed in the spring of 1980 by NIOSH to identify and evaluate articles that contained information on lung cancer mortality risks at and below the present permissible exposure limit. The conclusion drawn from their evaluation was that a two-fold excess risk of lung cancer mortality at and below 120 cumulative working level months (CWLM) of exposure to radon daughters is evident. This composite indication from selected studies was of sufficient magnitude to justify further evaluation. The study group, however, recognized that other studies and information must also be considered in a quantitative risk assessment which would form the basis for recommending an acceptable and feasible permissible exposure limit. A second factor in the decision to pursue further evaluation was the gaps in the current standard that had previously been identified by the Mine Safety and Health Administration (MSHA). These included lack of medical monitoring of underground miners and absence of regulations for surface miners. The seriousness of the health hazard relative to other hazards in mining was also considered. Along with exposures to silica and asbestos fibers, radiation was judged to be one of the major health hazards in mining, These combined factors constituted the justification to develop criteria and recommendations for improved mandatory health standards. The development process begins with a survey and review of the available world-wide information on the topic, including data and information developed by NIOSH. The end product of this review is to be a document that will contain an evaluation of the collected information and support for any recommendations that are made. To develop this document, the present work group has been divided into five task groups with the following emphases: Health Effects, Medical Aspects, Monitoring, Environmental Exposures, and Engineering Controls and Work Practices. The Health Effects task group is to evaluate the evidence from epidemiologic and animal studies of adverse health effects associated with all forms of ionizing radiation encountered in mining and milling operations. The "weight-of-evidence" of the results of all relevant and useful studies will be summarized, with the reasons for emphasizing the cited studies. The critical cells or tissues and the factors that should be considered in estimating the dose to these areas from various types of radiation will be identified. An evaluation will be made of the possible impacts of smoking and exposure to diesel exhaust on the determination of health effects related to radiation. The implications of biologically redundant dose in terms of the time between tumor initiation and death will also be analyzed. The Medical Aspects task group is to review the generally-accepted and the state-of-the-art medical technology for the detection of adverse health effects from ionizing radiation exposure. An evaluation will be made of the accuracy of urine and fecal analysis, wholebody counting, chromosome analysis and nose blows, as well as their usefulness for early detection of adverse health effects. Early detection is of little utility to the affected individual unless subsequent medical care can improve the prognosis. Recommendations for screening tests will be made after carefully considering the accuracy of the diagnostic procedures and the usefulness of early detection. Required recordkeeping and transfer rights will also be addressed. The Monitoring task group is to consider the state-of-the-art technology for the monitoring of occupational radiation exposures. Instrumentation, sampling strategies and analytical procedures will be reviewed for both personal and area sampling. The implications of the associated levels of confidence for non-compliance decisions will be evaluated. The discussion will also include an evaluation of the feasibility of replacing present monitoring systems with recent technology and the impact that a possible lower permissible exposure limit would have on monitoring requirements. The Environmental Exposure task group is to investigate the field procedures and mathematical methodologies which have been used to quantitate exposure levels to the various kinds of ionizing radiation. The magnitude and direction of the possible biases in past exposure assessments will be estimated. Specific attention will be given to the controversies concerning the quantification of the biological dose equivalent and the feasibility of recommending standards for mixed radiation exposures which use "rem" as the unit of measurement. The Engineering Controls and Work Practices task group will analyze the advantages and limitations of

Jan 1, 1981

By S. Horton, F. L. Stanley, P. J. Bethell

Introduction The McClure River Processing Plant is located at McClure in Dickenson County, VA. The plant was originally constructed in 1979 as a 600 st per hour facility. The plant is fed by coals from numerous mines in the area as well as a major longwall mine onsite. Coal is processed predominantly for the metallurgical coal market. The original plant was comprised of three separate processing circuits incorporating a heavy medium vessel for the coarse material (+10 mm). Intermediate sized material (10 x 0.6 mm) was beneficiated in heavy medium cyclones, with the remaining fines (0.6 mm x 0) passing to froth flotation for upgrading. The products from all three circuits were combined prior to thermal drying and then stored in silos or ground storage prior to shipment to the metallurgical coal markets. A middlings heavy medium cyclone circuit was also present to rewash the vessel and primary heavy medium cyclone rejects. The original circuit is as shown as Fig. 1 (Editor's note: All figures are reproduced at the end of this paper.) In reasonably close proximity to the McClure Plant is the much older, less efficient Moss 1 Plant, (Chance cone, table, water only cyclone, froth flotation, Fig. 2), which prior to the McClure Plant expansion was processing considerable tonnages of coal. At that time, the McClure River Plant was processing coal at the maximum possible feed rate, and the Moss 1 Plant was finding increasing difficulty in maintaining the stringent quality specifications on its major metallurgical coal accounts. Given the fact that many of the feeder coals to Moss 1 could be moved with minimal difficulty and cost to McClure, an exercise was conducted to evaluate the potential of expanding the McClure Plant. This would result in transferring coals from Moss 1, thereby increasing yield, reducing operating costs, and enabling customer specifications to be met more easily. After evaluating many possible ways of expanding the McClure Plant, we eventually settled on an approach based on the idea of Gallagher at German Creek in Australia (Gallagher, 1986). Our objective was to increase plant capacity by removing the coarser size fractions (0.60 x 0.15 mm) from froth flotation and at the same time eliminating the finer fractions (1.25 x 0.6 mm) from the heavy medium cyclone circuit, thereby allowing these circuits and therefore the plant as a whole to increase its capacity. These size fractions (0.60 x 0.15 mm from froth flotation and 1.25 x 0.6 mm from the heavy medium cyclone circuit) would be treated in a new spiral circuit. After desired evaluation of the various circuit capacities, including desliming screen, clean coal filter and thermal dryer capacities, etc., it was concluded that we should be able to maintain high levels of processing efficiency in the existing circuitry while at the same time expanding plant feed rate from a nominal 600 st per hour to 1000 st per hour, thereby achieving the objectives The circuit that was finally settled on is shown in Fig. 2. Several unique features were incorporated in this circuit in an attempt to maximize processing efficiency and reduce maintenance costs. First, the idea of a single-stage spiral circuit was abandoned in favor of a two-stage circuit incorporating rewashing the primary spiral middlings in a secondary spiral (Bethell, 88). The spirals selected were Mineral Deposits Model LD4 units. It was felt, after much previous testing, that spiral middlings rather than being a true middling gravity fraction were in fact a combination of true middlings with a preponderance of misplaced light coal and high-density rock. Consequently the insertion of this material directly into either clean coal or refuse streams (per normal practice) would only lead to loss of efficiency. The operating philosophy behind the design of the circuit was to produce the purest coal and rejects possible by adjusting the spiral splitters. Misplacement of refuse into the clean coal and vice versa would be minimized by taking a fairly large middlings cut. By reprocessing this material (10%-15% of primary spiral feed) in a secondary bank of spirals the misplaced material in the primary middlings would be reprocessed and its probability of reporting to the correct stream greatly enhanced. To minimize fines loading on the spirals, fine aperture sieve bends were to be included. A further novel feature of the circuitry is the desliming/ dewatering system. A philosophy of accepting inefficiency but forcing it to work in our favor was developed. The dewatering screens (Linatex Combis) were chosen not for their ultimate screening efficiency but for their ability to remove slimes (tolerating a fair quantity of oversize in the effluent) and their enhanced maintenance characteristics. Any oversize passing into the effluent, together with the fine coal centrifuge effluent, would pass to classifying cyclones, the objective of which was to reject to overflow (and hence froth flotation) as many fines as possible (-150 µm) without running the risk of losing coarse (1.25 x 0.15 mm) material to oversize. Hence, by a process of producing a "slime free" screen product coupled with a "coarse free" cyclone overflow, sizing efficiency could be maximized. Obviously, recirculating load had to be factored into the Combi Screen design capacity. The spiral circuit as shown in Fig. 2 was installed at McClure River in January 1988. The spiral circuit efficiency was very high immediately, and, after the normal debugging, plant capacity has been increased to the point at which we are currently processing of the order of 1100 st per hour. Test program The performance of the spiral circuit has been evaluated by two major samplings of the entire new circuit, coupled with routine bi-monthly tests on all other major components of the plant circuitry.

Jan 1, 1992

A Directory of Computer Programs Applicable to US Mining Practices and Problems. A. L. Sanford, T. L. Myers, and J. F. Stiehr. USBM OFR 131-78, Nov. 1977. A Master Environmental Control and Mine System Design Simulator for Underground Coal Mining. Tape MDS-5. Pennsylvania State University. USBM CT 1.77, Dec. 1975. Contains computer programs on a dynamic general purpose computer simulation model for underground coal mining. A Method for Computing Stabilization Pressures for Excavations in Incompetent Rock. J. D. Dixon and M. A. Mahtab. USBM RI 8128, 1976. Describes technique for determining the confining pressures that must be provided for stabilizing underground openings in incompetent rock. Accident Cost Indicator Model to Estimate Costs to Industry and Society From Work-Related Injuries and Deaths in Underground Coal Mining. Vol. I.III. D. G. DiCanio, A. H. Nakata, D. Colvert, and E. G. LaVeque. OFR 39(1)(3)-77, Dec. 1976. Describes a computer-based model for estimating the tangible costs of injuries and deaths from work-related accidents In underground coal mines. An Interactive Computer System for Evaluating Coal Mine Illumination. R. Goldstein. USBM OFR 110-80, March 1980. Computer system for calculating the illumination on coal mine surfaces due to machine-mounted lights. Analytical Modeling of Mine Roof Behavior Using Statistical Material Properties. R. T. Langland and F. J. Tokarz. USBM UCRL-51876, Aug. 1975. Analytical study of coal mine roof behavior using a computer model of a room-and-pillar mining method. Application of a Total System Surface Mine Simulator to Coal Stripping. Vol. I-VII. Pennsylvania State University. USBM OFR 33(1)(7)-78, Aug. 1978. Describes model development, input and output documentation, three case studies of an open-pit materials handling simulator, and development of two equipment selection models and a cost model. Characterization of the Structural Behavior of Rock Masses. Vol. I-II. L. R. Herrmann and M. A. Taylor. USBM OFR 67(1X2)-75, Sept. 1974. Computer subroutines for characterization of rock behavior, Including variability of properties within the rock mass, planes of weakness orthotropy, strength criteria, and post-failure behavior. Coal Mine Electrical System Evaluation. Vol. I-VII, L. A. Morley, et al. USBM OFF! 61(1)(7)-78, 1977. A concept to improve underground coal mine electrical system safety and availability. This proposed technique is based upon the ability to predict incipient failures in the mine power system. Prediction is made possible by a new method which uses a minicomputer to implement pattern recognition algorithms. Computer Modeling of Fluid Flow During Production and Environmental Restoration Phases of In Situ Uranium Leaching. R. D. Schmidt. USBM RI 8479, 1980. Describes development and application of a computer model for simulating the hydrological activity associated with in situ leaching. Interpretation of Rock Mechanics Data, Vol. II (A Guide to Using UTAH2). W. G. Pariseau. USBM OFR 47(2)-80, June 1978. Brief description of a two-dimensional elastic-plastic finite element program intended for mine stability analysis. Iterative Approximation Techniques for Microseismic Source Location. G. H. Dechman and Meng-Cherng Sun. USBM RI 8254, 1977. Iterative approximation method for microseismic source location developed to adapt microseismic techniques for rock structure analysis to investigations of surface problems such as slope failure in an open-pit mine. Limits and Cost Sensitivity of Alternative Parting Handling Methods. Vol. I-II. T. E. Finch, D. R. Haley, and C. J. Speaks, Jr. USBM OFF! 117(1)(2)-77, March 1977. Programs to examine alternative methods of handling the parting between two coal seams in a surface mining operation. Alternative methods consist of equipment implementation in either a stowing or haulback technique. Methane-Water Flow in Coalbeds. H. S. Price. USBM CT 1-79, April 1972. This computer program calculates transient, two-dimensional, gas¬water flow in a porous medium. Plastic Canopy-A Computer Program for the Structural Analysis of Protective Canopies. K. D. Winters, G. A. Gavan, and J. C. Ault. USBM IC 8795, 1979. Presents a FORTRAN IV program for the structural analysis of protective canopies used In underground mines and rollover protective structures used in surface mines. The Three-Dimensional Structural Analysis of Double-Entry and Single-Entry Coal Mines. Vol. I-III. Agbabia Associates. USBM OFF 16(1)(3)-80, Oct. 1978. Describes the computer program designed to perform a three-dimensional finite element analysis of an instrumented section of double- and single-entry coal mines. Underground Mine Evacuation Plans Analyzed by Computers. D. Tesarik, D. Nicholson, and A. B. Boghani. MSHA, Vol. 4, No. 5, Oct. 1979. USBM has developed a computer program to evaluate and construct mine evacuation procedures and plans.

Jan 11, 1981

By V. B. Achanti, A. W. Khair

Introduction Laboratory studies were carried out at West Virginia University to investigate the rock-fragmentation mechanism of continuous miners using an automated rotary cutting simulator. The primary factors influencing the fragmentation process were found to be bit spacing, bit geometry, depth of cut and cutting-drum rotational speed. This paper presents a discussion of the effects of these parameters in achieving optimum energy consumption and minimizing dust generation during rock fragmentation. The removal of rock ridges/walls between adjacent grooves is analyzed with three hits mounted simultaneously on the cutting head, while the bit tip angle was varied from 600 to 900. Bit spacing was varied from 25.4 to 50.8 mm (1 to 2 in.) while the cutting process was assessed for varying cutting depths. Respirable dusts generated during the course of the experiments were analyzed utilizing cascade impactors. Assessment of these parameters has led to a better understanding of the cutting mechanism of continuous miners in terms of energy consumption and dust generation. A review of the literature revealed that a considerable amount of research has been carried out on rockcutting processes. Many authors agree that the mechanical cutting efficiencies of mining machines (e.g., continuous miners, shearers and road headers) are affected by a host of parameters. Some of these parameters are machine controlled, some are operator controlled, while others are uncontrollable. Efforts were focused on understanding the influence of parameters such as bit spacing, cutting depth, attack angle, bit type, drum speed, bit geometry (i.e., tip size, shape and tip angle) and rock type on the cutting process efficiency in terms of specific energy consumption and respirable dust generation (Strehig?? et al., 1975, Hanson et al., 1979, Khair et al., 1989). Roepke et al. (1976) in an attempt to study the dust and energy generated during coal cutting using point attack bits found that the dust and the specific energy consumed both decrease as the depth of cut increases. The four fundamental stages of dust generation luring rock fragmentation are identified by Zipf and Bieniawski (1989). Coal breakage by various types of wedges was assessed by Evans and Pomeroy (1966) in an extensive experimental study on British coals. Yet the industry today requires further attention and guidance to optimize the energy consumption and dust generation during the rockbreakage process. This paper attempts to give a better understanding of the influence of some of the parameters listed above and focuses on further improvement in the rock-cutting process. The specific energy consumed for different types of bits used and the respirable dust generated are analyzed in the context of the variation of a few other parameters. Laboratory investigation The experiments were performed in the Rock Mechanics Laboratories located at West Virginia University. A rock-cutting simulator designed and fabricated by Khair (1984) was utilized for this purpose. The details of this machine are available in the literature (see Khair 1984). For this study, a series of preliminary experiments was carried out to determine the optimum cutting frame advance speed. This was intended to facilitate a maximum cut depth of 31.75 mm (1.25 in.) at an advance rate of 0.525 mm/s (0.0207 ips) for all types of bits being used and various bit spacings being considered. To look into the cutting-process efficiency of a continuous miner in the laboratory, several parameters of influence are being considered. Besides the bit-geometry parameters, machine- and operator-controlled parameters, such as spacing of bits on the cutting head, the cutting head rotational speed and the total cutting depth during an experiment, are varied. At the time this paper was written, only part of the completed experiments were ana¬lyzed, and a number of experiments were still being carried out following an orthogonal fractional factorial experimental plan to assess the effect of all of the above¬mentioned parameters on the cutting efficiency in terms of energy consumption and dust generation. Three different types of tip angles, namely, 60°, 75° and 90°, and two different tip sizes, namely, diameters of 7.94 and 24.61 mm (0.313 and 0.969 in.), were used. At

Jan 1, 1999

By Bond. F. C.

Introduction Although autogenous grinding applications date back to the early years of mineral processing when tumbling-type mills were first used, there has been a renewed interest in this type of grinding over the past few decades. A review of the historical aspects of autogenous grinding is provided which follows the development of the various types of mills and processes. Wet and dry grinding operations using either autogenous or semiautogenous mills, including the Aerofall mill, are described. Em¬phasis has been placed on the test work required to evaluate the feasibility of using autogenous or semiautogenous grinding for a spe¬cific application. The factors which must be considered to establish the appropriate design criteria for these types of grinding circuits are discussed in detail. Successful application of these grinding circuits results from a thorough test program, a careful analysis of the test results, and a practical engineering approach to the design of the process flowsheet and milling equipment. Certain pebble milling applications, where the pebbles consist of the ore or rock itself, can be classified as autogenous pebble mill grinding, or better known as ore-pebble mill grinding. In the generic sense, pebble milling includes those grinding operations using a tum¬bling mill in which the grinding media consists of various types of pebbles, either ore or other pebbles, instead of metal balls. Conse¬quently, this subject is covered separately in Chapter 5, Pebble Mills. Operating data from many of the installations located throughout the world are included in this chapter. Our appreciation is extended to those companies who contributed this important information. History of Autogenous Grinding F. C. BOND Autogenous grinding, the grinding of ore by itself rather than by special metallic or nonmetallic grinding bodies distinct from the ore, has an old and honorable place in the art of mineral processing. It antedates the use of the rod mill and is only a few years younger than the development of the continuously fed rotating tumbling mill in which it is accomplished. Its history71, 72 is inextricably bound up with that of gold mining. Every important early development of autogenous grinding took place on ores of gold, except in one case of silver ore. In common with many other features of modern mineral processing, autogenous grinding would have been greatly retarded if the winning of gold had been so small a part of mining in the past as it is today. Gold unlocked the secrets of rock-on-rock grinding. In years past gold ores were crushed in stamp batteries and then flowed over copper plates covered with mercury which amalgamated the gold. The McArthur Forrest cyanide process for the dissolution of gold in a dilute alkaline cyanide solution appeared in 1890. With finer grinding it recovered more gold than could be won by amalgama¬tion alone. The first response was to stamp in cyanide solution and finer mesh screens guarding the stamp discharge. However, the fine screens greatly restricted the tonnage stamped, and they wore out rapidly. Rotating tumbling mills70, 73 were the obvious choice to grind the coarse product from stamps. Many of these were Gates tube mills; more than 1,000 of these were sold between 1907 and 1913. They were made in 6, 51h, and 5 ft diam, with lengths of 20 and 22 ft. The 5 x 22-ft size was the most popular. The term tube mill did not refer to the mill shape, but included all tumbling mills using grinding pebbles. The flint or chert rocks found along the coasts of Denmark and Normandy are probably the most durable and resistant to abrasive and tumbling wear of any on earth. They were much used in the early tube mills. The early linings were all silex or Danish flint blocks, cemented into place with portland cement. Local hard stones or cast iron blocks were also used later. The cast iron liners carried lifters and were not bolted to the shell, but were held in a tight circle against it by steel wedges hammered home between them. The first ribbed metal liners designed to hold grinding pebbles pounded into them by the mill action were developed at El Oro, Mexico, about 1905. These and the Forbes, Komata, and Tonopah types, as well as the Osborne type developed in South Africa within about five years, were very successful in reducing liner wear. They sometimes lasted two years or more, and equalled or exceeded the life of silex lining at a lower cost. Some of the types were patented. The pioneer name recorded in the history of autogenous grinding is that of Kenneth L. Graham. In 1907 at the Geldenhuis Deep Ltd. mine near Johannesburg, he ran a comparative test on two tube mills. One mill used pieces of the common gold ore, or banket, and the other used the customary imported Danish pebbles as grinding media. He published the first account of autogenous grinding.'' In a test lasting 81 days the ore pebbles showed a definite saving over the Danish pebbles. Graham's historic paper also describes the invention of B. Chew of the Crown Deep Mines of the first trammel to be attached to a grinding mill discharge. The first known account of autogenous grind¬ing printed outside of South Africa was the description of the Graham test in Mining and Scientific Press's The use of grate- vs. overflow¬discharge was also described at this time." The news of this successful test appears to have had immediate effect, for by the end of 1908 the number of tube mills on the Rand had increased from 72 to 120.76 Presumably, most of these mills were now using autogenous grinding. The ore was principally a hard conglomerate of quartz peb¬bles cemented by silica, and was eminently suitable for use as a grind¬ing media. It is unfortunate that in South Africa the new autogenous process continued to be called by its old name of pebble milling, since the name does not distinguish between grinding with ore and grinding with extraneous pebbles. For increased clarity the former is called rock pebble milling. The first historical stage in the development was what is now designated as secondary autogenous grinding, or rock pebble milling. It consists of grinding a feed all passing 'A in. or finer with ore pebbles about 2 to 4 in. in diameter. The next stage to develop was that now called intermediate autogenous grinding, where larger amounts of larger pebbles about 5 in. in size were used to grind ore which had been crushed to about rod mill feed size or somewhat finer. It was developed on the Rand and in other places where coarse stamp screens were used to increase mill tonnage. In many cases the early secondary and intermediate autogenous mills have now been replaced by conventional ball mills and rod mill-ball mill circuits, with the object of further increasing grinding capacity. The third stage to develop is designated as wet primary autogenous grinding, in which run-of-mine ore or a primary crusher product is all fed to the autogenous mill together. It developed many years later indepen¬dently in the United States and in South Africa. The fourth and most recent stage is that of dry primary autogenous grinding, which developed in the United States and Canada. All four stages originated in gold mining. The first use of commercial autogenous grinding outside of the Rand apparently took place five years later in 1912 at the Santa Gertrudis gold mine at Pachuca, 55 miles northeast of Mexico City. This was followed closely by an installation at Goldfields, Nevada, in 1914-1915. The first name associated with autogenous grinding

Jan 1, 1985

By Donald C. Moore

Introduction Asamera Minerals (US) Inc. and its joint venture partner, Breakwater Resources Ltd., discovered ore grade gold mineralization on their 20 km2 (5000 acre) Wenatchee, WA land position in February 1983. Due to the high grade nature of the discovery ore and the known reserves of ore in the "B Reef' and "B West" zones previously outlined by other companies, a decision was made to construct a mine/mill operation near the known ore occurrences. Further drilling in the discovery area quickly expanded known gold occurrences to more than 3.6 Mt (4 million st) with tentative in-place ore grade of 7 g/t (0.25 oz per st) and minor silver values. Based on existing knowledge of the ore body and the rapidly increasing ore reserve, a decision to build a 1.8-kt/d (2000-stpd) mine and mill complex was made in the second quarter of 1983. A schedule was devised to begin immediate mine development, shaft sinking, environmental and land use permitting, and mill and tailings dam construction (Fig. 1). Meeting the scheduled startup date, April 1, 1985, required a fast track schedule in all areas. To this end, Asamera purchased the Oracle Ridge Partners concentrator. This was an assemblage of new equipment designed for use as a copper concentrator in southern Arizona. The purchase contained all of the major mineral dressing equipment - crushers, screens, rod and ball mills, etc. and an engineering package. It did not include most of the other required items, such as buildings, conveyors, pipelines, tanks, and pumps. At the same time, core samples were sent to two independent process development laboratories for initial flowsheet development. Due to the refractory nature of the carbonaceous ore, cyanide leaching was not feasible. Flotation was selected as the concentration process. Further testing showed that autoclaving of the flotation concentrate followed by cyanidation would result in overall recovery of about 85% gold. A mine manager was hired to begin assembling an operations staff, hire an environmental consulting firm, and begin mine development. Environmental and land use concerns were major obstacles due to the mine's close proximity to a city of 20,000 people. These concerns had to be rapidly defined so as to mitigate any adverse impacts from and mining processing operations. Baseline data dealing with weather, air and water quality, and sound were measured before start of mine construction. Concentrator and flowsheet development remained static until October 1983 while definition drilling and mine development proceeded. In late October, a process engineer was hired to coordinate development of a process flowsheet, purchase the remainder of the concentrator equipment, prepare a concentrator construction contract, finalize concentrator detail engineering, and combine environmental and process requirements with a tailings dam design. Process development There were only 17 months remaining to mill start up from the hiring date of the process engineer. Therefore, the process flowsheet had to be finalized rapidly. To accomplish this, samples of drill core from the highest grade (and therefore potentially the most commercial) ore zones were sent to an outside metallurgical laboratory to confirm beneficiation tests on the flotation process. Test results again showed that flotation would provide about an 86% gold recovery. Therefore, all further testing was concentrated on flotation and autoclave/cyanidation of flotation concentrates. Focusing on a well known process such as flotation was important in accomplishing the rapid design and construction of the concentrator. If, during these next phases, we were continually changing design concepts, layout, and process flow, the mill startup would have been delayed many months. Once a process flowsheet is selected the process engineer must obtain the process criteria needed to design the beneficiation system. For example, it was known in early December that the Oracle Ridge rod and ball mills were too small to grind 1.8 kt/d (2000 stpd) of Wenatchee ore. A decision had to be made to purchase a large, used ball mill and convert the Oracle Ridge ball mill to a rod mill. The process engineer must be cognizant of the process criteria needed to size and select equipment. If not, the process engineer must use the professional services of the equipment manufacturing companies to review the requirements that the equipment is asked to perform. For the Wenatchee system, this resulted in the adaptation of a ball mill to a rod mill with a weight limit of grinding rods to protect the mill bearings and drive trains. When a decision is required, the process engineer has to present the facts and options in a manner that allows a rapid decision. This information must include costs, equipment availability, and effect on the construction schedule. At the Cannon mine, there were process development details that resulted in decisions similar to the ball mill purchase. These included an increased flotation residence time from eight to 25 minutes, an increased thickener area requirement, a high pressure tailings pumping system, and area constraints in plant layout. All of these decisions had to be timely and required assistance from manufacturers' service engineers, and knowledge of the alternate costs and effects on construction completion. Equipment procurement It was decided in early 1983 to build the ore milling facility with Oracle Ridge equipment, augmenting it with used equipment

Jan 2, 1989

By P. Wearden, K. Bryner, K. Vrana, V. Castranova, R. Dey, R. Reist, J. Blackford

INTRODUCTION Proliferation and enhanced synthesis of collagen by pulmonary fibroblasts have been shown to be key steps in the development of chronic silicosis (Goldstein and Fine, 1986). The regulation of lung fibroblast proliferation by cytokines released from alveolar macrophages may be an important pathogenetic mechanism in the development of the fibrotic process (Kelley, 1990). One cytokine, platelet-derived growth factor (PDGF), promotes fibroblast proliferation by inducing the movement of quiescent (Go) cells into the C1 phase of the cell cycle (Chen and Rabinovitch, 1989). Others regulate the rate of transition of fibroblasts from Gl into the S phase (Leof et al., 1982). These two classes of cytokines have been termed, respectively, competence and progression factors. One approach used to examine the release of cytokines from macrophages is the fibroblast proliferation assay in which fibroblasts are exposed to culture supernatants from macrophages exposed to various stimuli. In most of these assays, the supernatant contains fetal calf serum which provides the competence factor(s) necessary to facilitate the proliferation of fibroblasts (Bitterman et al., 1982; Bitterman et al., 1983; Elias et al., 1988). Recently, a fibroblast proliferation assay using plateletpoor plasma (lacking competence factor(s)) as a substitute for fetal calf serum has been described (Kuman et al., 1988; Bauman et al., 1990). In this assay, the release of a competence-inducing PDGF-like growth factor from rat and human macrophages can be distinguished from other cytokines that act as progression factors. In order to obtain more consistent results and with the ultimate goal to be able to discriminate between the effects of competence factors as opposed to progression factors, we have conducted experiments to determine the appropriate concentrations of plasma and PDGF required for imparting competence in the fibroblast proliferation assay. We tested lung fibroblast cells obtained from explants of rat lung tissue and also a fetal human lung fibroblast cell line obtained from American Tissue Culture Collection (ATCC153). MATERIALS AND METHODS Fibroblasts Specific pathogen-free, male Sprague-Dawley rats were use in some studies. Animals were given a lethal intraperitoneal dose of sodium pentobarbital. Fibroblasts were isolated by chopping the lung in enzymes that digest the connective tissue but liberate lung cells for further study (Rabovsky et al., 1989). After digestion, the remaining lung tissue suspension was filtered through two layers of sterile gauze and centrifuged to recover lung fibroblasts. These were resuspended in culture medium that contained 10% fetal calf serum and distributed to culture plates for growth. In other experiments, a human fetal lung fibroblast cell line, obtained from American Type Culture Collection, Rockville, MD, 20852, was used instead of rat lung fibroblasts. In these cases, a 1 ml ampule containing human fetal fibroblasts was plated into a tissue culture flask containing medium plus 10% fetal calf serum. For both types of fibroblasts, culture medium was changed 3 times per week and cultures were incubated at 37°C until confluent. Harvested rat and human lung fibroblasts were quantified using an electronic cell counter equipped with a cell sizing attachment (Coulter Electronics, Inc., Hialeah, Florida). Tritiated Thymidine Incorporation The basic procedural outline of Kumar et al. (1988) was used with modifications to evaluate tritiated thymidine incorporation into fibroblast DNA following exposure to PDGF and plasma. Both rat and human lung fibroblasts were plated at 50,000 cells/ml at a density of 250,000 cells/25cm2 culture plate. Cells were quiesced for 4 days with 2% rat plasma. As the assay was refined, fibroblasts were quiesced in plasma-free media for 48 hrs, since the mitogenic activity of 2% plasma was variable. Test medium was applied for a period of 6 hrs, followed by a 24 hr tritiated thymidine (lµCi/ml) labelling period in plasma-free media. Medium alone was used as a negative control and media with 10 or 20% fetal calf serum was used as the positive control for rat and human fibroblasts, respectively. Cell Quantification and Measurement of Mitogenesis Twenty-four hours after the addition of tritiated thymidine, the fibroblasts were washed with 5ml of fresh serum-free media, centrifuged and resuspended in phosphate-buffered saline. The cells were dissolved in 0.5m1 of O.1N NaOH and radioactivity determined in a beta counter. Incorporation of trititaed thymidine as an index of DNA synthesis was expressed as DPM/fibroblast. RESULTS In the present study, we quantified mitogenic potential by monitoring the incorporation of tritiated thymidine as

Jan 1, 1991

By John P. Cummings

Glass is the most unique substance which man has made. It is understood that Phoenician sailors playing on the beach in Mesopotanmia around 12000 BC, produced glass by the fusion of sand and soda in their campfire. Glass is unique for many reasons. However, let us consider the following. It is made of materials which are opaque and yet glass can be translucent, transparent, or opaque. It can be as soft as cotton or stronger than steel. It transmits light readily and yet is totally impervious to gases and liquids. It is chemically an inert, creates very little pollution in its manufacture. It is made of the most abundant raw materials on earth. It is readily recyclable and has little pollution in its disposal. Glass containers are made principally from sand (that is, silica), soda, which is sodium carbonate, and lime, which is calcium carbonate, plus decolorizers or colorants and cullet. The proportion of cullet, broken glass, in most glass tanks today is roughly 10 to 20% of the glass product. In most cases, the cullet used in the plant is glass from off-ware, or broken ware, which is a waste product within the plant. Cullet has the following advantages. It contains all of the raw materials normally used in a batch. It tends to melt at a lower temperature than batch, thus saving some fuel, and, finally, by forming a crust above the normal melting system, there is a tendency to cut down air pollution.

Jan 1, 1975

By K. T. Mau

Hawaii is located approximately 2,500 miles from the mainland, but has basic needs for its thriving developing community. Cement is one of these needs Cement production in Hawaii is not without handicaps. The necessary raw materials are practically non-existent except for limestone. Even at that, the available limestone is marine in origin, coralline in nature and amorphous in characteristics - unlike calcite which is normally the calcium carbonate source for cement manufacture. There is no clay or shale for the argillaceous portion of cement raw mix. Hawaii being volcanic in origin is too young geologically for clays and shales to evolve and whatever there is, most likely is washed into the ocean.

Jan 1, 1982

By S. A. Braley

Water is our most important natural resource, for without it there is no life. It is also the principal natural resource for all manufacturing processes, in particular for the development of power. For these reasons it is essential that water be protected and conserved, with the realization that its quality 18 as important as the water itself. A significant factor in the quality of ground water is the mining of coal -- that other main ingredient of Power. In the mining of coal it is necessary to disturb the geological strata overlying or adjacent to the coal seam. This procedure not only releases water from water-bearing strata or underground aquifers but also exposes other minerals which eventually become the source of pollution or deterioration of the released water as well as of the surface stream into which It is discharged It is then evident that in attempting to utilize one essential natural resource there is degradation of the quality of the other. Many selfishly assume that the principal function of a stream is fishing and recreation, but the real function is the support of industry and agriculture upon which civilization is dependent.

Jan 1, 1959

By Miroslav Cervenka

The Coal, characterized by coal quality parameters as for example, ash content (Ad), moisture (Wrt), heat content (Qr1), sulphur content (Sd) and other parameters is still in many countries the basic source of energy. Energy is life, especially in modern society, where the demand for ecologically clean energy is very high; and therefore, there is great importance in effective coal utilization Direct utilization - burning of coal for energy generation, is still one of the most effective methods. In most cases, it is one of the cheapest energy transformation methods in spite of the fact that this energy transformation is connected with the deterioration of our surroundings. Deterioration can be substantially reduced if the burning of coal at power stations, heat power stations and other coal consumers is optimized. That means the burning is done in steam boilers, the technical state of which is very good; the coal has minimimum .variations with respect to the projected values; and the burning process is controlled and checked not only from a regulation point-of-view but also from an ecological point-of-view. The instrumentation capable of measuring the characteristic coal quality parameters in real time exists and it forms the basic assumption for the work our research team has made. The proper application of this instrumentation is able to provide information for burning process optimalization. It is the main topic of this paper, which is the compilation of the information on our work -made during the last five years. The whole information is based on practical experience and applied systems in Czechoslovakia in the mines and power engineering industry. See [Cervenka, 1989]

Jan 1, 1992

By Randal J. Barnes

There are many methods available for the estimation of geologic parameters (Popoff, 1966; Barnes, 1980). In the last decade a new and relatively sophisticated technique using statistical theory has emerged: geostatistical kriging (David, 1977; Journal and Huijbregts, 1978; Rendu, 1978). An interesting property of geostatistical estimation is that when a geologic parameter is estimated for a given volume, not only is the estimate generated but an associated measure of certainty is produced as well. This statistical measure of estimation confidence is called the estimation variance. While the concept of an estimation variance is universally considered to be a good idea, mining engineers and geologists do not currently have the tools to incorporate it into their planning and quantitative analysis. This paper attempts to develop and present some of the necessary tools required to incorporate estimation uncertainty into mine planning. Geologic uncertainty is portrayed as an economic business risk. This risk is objectively characterized, and-the associated cost is quantified. Geologic delineation sampling is viewed in this light as an economic commodity with an associated cost and value. A coal mine planning example is used as a focus in this presentation. The data used is real, though it has been modified to maintain data security and project anonymity.

Jan 1, 1986

By L. G. Austin, R. S. C. Rogers

INTRODUCTION Coal is geologically a sedimentary rock with distinct bedding planes containing a mixture of petrographic constituents. Yet. because of its chemical nature, it also has some of the characteristics of a polymeric material. In addition, it is extensively precracked to give a macropore structure and also has a substantial internal volume of microporosity. Add to this the mineral constituents dispersed in the coal in a variety of forms and sizes. and it can be readily appreciated that coal is a complex and heterogeneous material. It has to date not been possible to predict the particle size distributions produced by breakage of coals from a priori application of the principles of fracture mechanics. However, there is a substantial body of data resulting from empirical observation of size distributions produced under various conditions, primarily from grinding mills. It is the function of this review to outline how this empirical information can be used to predict product size distributions for those systems which have been studied systematically, and to suggest future avenues of research using similar methodology. FRACTURE THEORY APPLIED TO COAL It is, of course, possible to devise experiments for slow crack propagation in a coal. However, the evidence is that coal breaks by brittle fracture in the mills and mining machines of industrial importance and therefore the theory of brittle fracture is sufficient to illustrate the important features of the fracture of coal. In order to produce size reduction a solid must be stressed by the application of force. The force can be a tensile (stretching), compressive, or shear force, and the three dimensional stress pattern in the solid represents the tensile, compressive, and shear forces per unit area acting in any direction through any point in the solid. As an illustration consider the case of a simple one-dimensional tensile stress: then stress is defined as a=F/A where F is the force. A is the cross-sectional area of the solid normal to F and strain. E, is defined as the fractional change in the length of the solid. so that E= x/L where x is the extension of an original length of L. Coal behaves as an elastic material over short times (although viscoelastic creep can occur if the stress is maintained over long times) and if it is not stretched too far the coal returns to its original shape when the stress is removed; otherwise catastrophic failure occurs and the coal fractures at a stress termed the tensile strength. The failure under stress of an elastic material is termed brittle fracture. Elastic material fails at small stain and the stress-stain relation to where brittle fracture occurs is the empirical Hooke's law. a= YE, where Y is Young's modulus. The work done on the solid to take the solid from zero external stress to a stressed state by slowly increasing F up to a final stress u is the integral of the stress times the strain. This reversible strain energy is stored in the solid and. using Hooke's law. is u /2Y per unit volume of solid. If the solid is immediately loaded to u the work done is [EAuL] which Is a2 /Y per unit volume, that is, twice the reversible work.

Jan 1, 1988

By Roger G. Haga, Richard P. Campbell

Introduction Project scheduling . . . . • What is it? • Why is it done? • Who does it? • How is it done? • When is it done? These are some of the questions that will be addressed in this paper. It is intended that by exploring these questions, an understanding of project scheduling, its purpose and benefits can be achieved. The purpose of this article is not to describe in detail "how to schedule", but rather to give an overview of the scheduling function on a project. A successful scheduling effort is dependent upon the people responsible for developing, implementing, and monitoring the schedule, and the tools and procedures that are used. These three elements are integral components of project scheduling. What is it? A schedule is the integration of network relationships, resource constraints and the calendar. The schedule identifies when the activities "will" be done, while the network states when they "must" be done. Networking techniques are the most widely accepted scheduling tool for complex projects. The benefit of the network is the simplicity in illustrating, presenting and analyzing the dependency relationships of a project. The logic network is the foundation of the schedule. When calendar dates are applied to the logic network, it becomes a schedule which provides the basis for evaluation of project status, forecasting the occurrence of events and controlling the project. The sequence of activities that requires the longest time to complete determines the overall duration of the schedule. This longest sequence of activities is called the "critical path". The "critical path method" (CPM) is used to describe this particular schedule network methodology. The two common conventions of CPM networks are the "arrow diagram met hod" (ADM) and the "precedence diagram method" (PIM). There are arguments in favor of each of these two methods, and usually the one that is used is a matter of personal preference, project directive or corporate procedure.

Jan 1, 1986

By M. Javier

"Mining is the main activity that supports civilization. Actual mining presents contradictory concept to the health of the planet. Its business model is linear, fragmented and characterized by inefficiencies in the utilization of water, energy and mineral resources which generate waste and pollution. Before mining begins, it is critical to determine the vulnerability of the ecosystem and after mining, it is critical to determine the depletion of nature. Attending to these vital matters will require quantifying and measuring them. The purpose of this study is to provide and establish the metric to assess the performance and manage a sustainable future for our existence and that of our planet. This index is defined to measure and control Nature Friendly Sustainability (NFS) based on understanding nature. This index will determine the condition of nature constructed on the measurement of physical, chemical and biological equilibrium. Proposing a new mindset for mining design to satisfy nature’s needs would be to offer the same opportunities to future generations. Hence, sustainability is synonymous with equilibrium under a limited ecosystem. Nature is interconnected. Finally, this index of NFS is readable, standardized and relevant to keep nature’s needs in equilibrium. Establishing and NFS index is the next metric for humanity. INTRODUCTION There are a multitude of definitions of what sustainability is. Most of them are defined in terms of the corporate entity from which they emanate. The classical definition of management states that, “if we do not measure it, we cannot manage it.” Management of ecosystems is the definition of sustainability. Indicators of quantity and quality of sustainability are vital in order to create a metric for sustainability. These indicators will measure the depletion and degradation of nature as they occur through human intervention. The first step is to create a metric of sustainability to simplify its definition. The definition of sustainability is based on whether or not you consider nature to be finite or infinite, which is described in Nature Friendly Sustainability (NFS) [1]. The definition of NFS applies to any anthropogenic activity with an understanding of the equivalent activity in nature. That is to say, human activity with nature requires balance and equilibrium in order to be sustainable. The importance of this equilibrium is quantified in a set of metrics in order to establish measurement and acknowledge indicators for our dependence on nature. Sustainability is synonymous with equilibrium in any ecosystem that is limited and fragile. In the graph is the explanation of sustainability definition. See Graph 1"

Jan 1, 2017

By M. Gershon

The article is an excellent one, providing a basis for the conduct of the major tasks in underground mine planning and scheduling. I object to only one word. Since that word is featured in the title, it is a strong objection. The word used erroneously is optimization. There is no optimization performed in the article. By definition, an optimization is an approach that yields the best of all possible solutions. We can limit this definition to its narrowest sense, the optimization techniques of operations research such as linear programming, or broaden it to include less sophisticated search procedures that find the best mine plan or schedule. No such claim is made in the article. No optimization of any kind is performed. The major tool used is a simulation, which, in truth, is probably more appropriate for this problem than any optimization. But simulation does not optimize. It only allows for quick evaluation of many options, leading to the selection of the best plan "of those tried," not "of all possible plans." This article, of course, is not the first mining paper to make this error. It seems that the word optimization is used more to impress people than to refer to any act of optimizing. It is used in a way that implies "try to optimize" or "come close to optimize," but not to really do it. It is usually appropriate to quietly ignore this error. But when Dr. Chatterjee, who has done so much good work in true optimization projects and whose work I have come to admire, makes such an error, it is time to publicly dispute it. Having discussed this very point with him in the past, I am sure that he will agree. If he can support this position in his response, perhaps people will start to listen and we can reduce the number of false claims of optimization. This should be easy for him in this case since the approach he used, simulation, is the more appropriate one anyway.

Jan 1, 1987