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By Michael E. Mcllwain, Lee O. Nelson, Robert D. Rogers, James H. Wolfram

Examination of the economics and environmental consequences associated with phosphate fertilizer production documents the necessity for development of innovative production technologies. Evidence of phosphate ore's susceptibility to microbial alteration indicates that it will be an ideal candidate for industrial-scale bioseparation. There are at least two applications for the biosolu- bilization technology: (1) an industrial-scale bio- process similar to the presently used wet acid process and (2) production of small prills (bio- bead) containing unprocessed rock phosphate and the biological ingredients necessary to promote phosphate solubilization. At the present stage of development, the biobead technology is the preferred application. It is an energy-efficient "green" technology that will dramatically reduce the amount of solid waste and air emissions produced by the phosphate industry.

Jan 1, 1999

By J. A. Herbst, Z. Rogovin

In order to develop more accurate procedures for scale-up and design of semi-autogenous grinding mills through detailed mathematical models, a better understanding of the charge motion and rocklball segregation is needed. To achieve this a 0.61 m (2.0 ft) diam by 0.305m (1 .0ft) length plexiglas mill was constructed as a scale model of a commercial SAG mill. Rock motion in the rotating mill was simulated using various sizes of ceramic balls. Quantitative measurements of dynamic angle of repose, water and slurry motion, holdup and size segregation of the charge were taken. In addition, a video presentation (available to the public) was prepared to elucidate charge and water motion at various mill speeds, charge volumes, rocksize distribution, and flow rates.

Jan 1, 1990

By J. Chappell

The Nuggety Gully prospect is located 170 km (106 miles) south- east of Port Hedland and 40 km (25 miles) northeast of Marble Bar in the Pilbara Region, Western Australia. The area is held under title by Esso Exploration and Production Australia, Inc. The prospect covers a significant, but currently uneconomic body of molybdenum and copper mineralization of porphyry style, associated with Archean high level acid intrusives intruding sediments and volcanics of a narrow east-west Archean greenstone belt. The Nuggety Gully intrusive rocks have been dated at 3.2 b.y. ± 117 (de Later, oral communication). The current resource estimate for the deposit is 102 Mt (1 12 million st) of 0.1 1 % Mo and 0.15% Cu.

Jan 1, 1991

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

6.1 PRINCIPLES OF PLANNING GROUTING OPERATIONS FOR SHAFTS, DRIFTS AND TUNNELS The planning of integrated grouting of vertical, inclined or horizontal underground workings in saturated, fractured rock includes the following phases: 1. Planning for the collection and analysis of geological information concerning the characteristics of the struc¬tural deformation(s) that produced the details of the frac¬ture system being studied. 2. Planning for the acquisition of initial hydrogeological information produced by the flowmetric and hydrody¬namic investigations (described in Chapters 2 and 3). 3. Computing the initial engineering calculations of the ex¬pected or planned dimensions of the isolation curtains for each water-producing stratum (fracture zone) and the expected areal spreading pattern of the grout in each of these water-producing horizons. The selection of the lo¬cations, the initial optimal number of holes and the vol¬ume of grout needed to create each isolation curtain also are included in this phase. 4. The selection or development of a grout formula for the specific hydrogeological conditions that exist at the site. 5. Identification of a resource (deposit) required to produce the grouting clay. The planning of grouting operations must be supported by the acquisition of all necessary initial data, which in¬cludes: 1. The geographic location and administrative responsibil¬ity (ownership and regulatory constraints) of the mineral deposit and any peculiar features of the area where ex¬cavation of the underground workings is planned. The surface relief, the hydrographic network of the region, and the local climatic conditions must be evaluated prior to planning for buildings and waste discharges if any. 2. The economic status of the region, the available trans¬portation, the nature of the energy base, the water and heat supply sources, the location of construction mate¬rials suitable for making the grout all must be deter¬mined. 3. The geological and hydrogeological characterization of the work site must be accomplished. 4. The technical features of the shaft(s) and underground works must be characterized. 5. The organization's principal initial design plan features and the construction sequence of all buildings and struc¬tures must be developed. These must be correlated with the organization and schedule for excavating the under¬ground works. The technical characterization of a shaft or of other un¬derground workings must contain information about the purpose of the workings, the design depth, the planned excavation procedures along with cross-sections, and the parameters of the lining, if a lining is required. The geology of the section to be penetrated must include geological data obtained from existing exploration activities at the mineral deposit and conclusions about the shaft ex¬cavation conditions obtained from drilling technical moni¬toring holes and exploration drillholes. The interpreted stratigraphic regime, including the thickness, strike and dip of each stratum, must be shown in cross-section. The tec¬tonics, including all geologic structures, must be presented by cross-sections. The bedding orientations must be used to identify anticlines, synclines, monoclines and faults. The lithologic composition and the drillability of the rock com¬prising the section must be examined. The character of the rock's fracture pattern must be presented. Finally the per¬meability, porosity and storage characteristics of each aqui¬fer must be determined by in situ hydraulic property testing (injection tests or pumping tests, as described in Chapter 3). The fracture pattern of the indurated rock is presented in reports derived from geo-exploration borehole studies and from studies of outcrops on the surface and in all nearby underground workings. The extent of each fracture zone is mapped in three dimensions. The distribution of fracture apertures, the nature of the fractures according to each litho¬logic stratum, and the relationship of bedding planes to each fracture system are all determined. Acoustic logs, downhole cameras, caliper logs, oriented core, and any other appro¬priate geophysical tools available are used to construct the data base for this purpose. The hydrogeological characterization of the geologic section to be penetrated by the workings is based on the analysis of hydrogeological data collected during the drill¬ing and testing of technical monitoring holes and from in¬formation on ground water inflow collected during the ex¬cavation of nearby underground workings, if such workings exist. The hydrogeologic characterization must contain the following information: 1. The identification of all aquifers, their thicknesses and their locations in the section, 2. The distribution of hydrostatic head in the ground water in three dimensions, 3. The coefficient of saturated hydraulic conductivity (per¬ meability) of each aquifer, 4. The porosity and storativity of each aquifer, and 5. The projected rate of water inflow into the shaft, drift or tunnel for each aquifer prior to grouting. This step is used to identify the zones to be grouted.

Jan 1, 1993

By J. M. Smith, T. R. Horton, R. L. Blanchard, T. W. Fowler

INTRODUCTION Uranium mining operations release radioactive materials into both air and water and generate large quantities of solid wastes containing low levels of radioactive materials. Solid wastes produced by mining operations remain on the surface at many inactive mining sites in the Western United States. These mining effluents may present a potential health and environmental hazard. Therefore, Congress, in Section 114(c) of the Uranium Mill Tailings Radiation Control Act of 1978, instructed the Administrator of EPA to prepare a report identifying the location and potential health, safety, and environmental hazards of uranium mine wastes and to recommend a program to eliminate these hazards. Several facts and limitations helped shape the method and approach of the EPA study. Little information on uranium mines was available; measurement information that was available on uranium mine wastes was frequently influenced (biased) by nearby uranium mills; time and resources did not permit comprehensive field studies to provide additional data; and there are inherent variations between uranium mines and sites that complicate generic assessments of mine wastes. To accommodate these facts, the EPA developed conceptual models of uranium mines and made health and environmental projections from them. The models were based upon available data from the literature, supplemented with information from discussions with persons inside and outside the EPA, and by doing several short-term field studies in Texas, New Mexico, and Wyoming. When necessary, conservative (maximizing) assumptions were employed. This paper presents a brief account of a part of the EPA study dealing with the potential health and environmental effects caused by active surface and underground uranium mines. Airborne contaminants are emphasized, although solid and liquid effluents are also included. Due to the limited space, only the methods and parameters used and the results of the assessments will be presented here. Anyone interested in the source of the data used and the development of the parameters should refer to the EPA report (Blanchard et al., 1981). The occurrence and emissions of stable elements were included in the EPA report, however, due to space limitations and their apparent small impact, except for possibly at some specific mines, only radioactive sources will be included in this presentation. MODEL URANIUM MINES The model surface mine was located in the South Powder River Basin of Wyoming and the model underground mine was located in the Ambrosia Lake area of New Mexico. These are the prevalent type mines in those areas. The model mines were based on the average production parameters of the 63 open pit mines and the 256 underground mines that were operating in the United States in 1978 (Department of Energy, 1979) and on a report of an extensive study of open pit mines in Wyoming (Nielson et al., 1979). Information contained in environmental impact statements and in reports from federal and state agencies was also used. Parameters for the model mines are listed in Table 1. The surface mining scenario is that 7 pits are opened in the 17-year mine life with overburden from each successively mined pit used to backfill a previously completed pit, resulting in an equivalent of one pit of overburden (2.4 year production) stored on the surface. No backfilling is assumed at the underground mine. Overburden or waste rock, ore, and sub-ore are separated into separate piles that are either rectangular in shape with length twice the width or in the shape of a frustum of a regular cone. Both shapes have 45 degree sloping sides. To account for bulking, the volume of the material comprising the piles was considered to be 25% greater than the volume of material removed from the ground. It was assumed that dewatering was required at both mine sites. Wastewater discharge rates at the surface and underground mines were assumed to be 3.0 and 2.0 cu m per min, respectively. SOURCE TERMS The following radioactive contaminants at active uranium mines were assessed in the EPA report: 1. Radioactive particulates in a) wind suspended dust from waste rock (overburden) pile, sub-ore pile, ore stockpile, b) suspended dust from mining activities (rock breakage, loading and unloading ore and wastes), and c) vehicular dust, 2. Rn-222 emanation from waste rock (overburden) pile, sub-ore pile, ore stockpile, and mining activities, 3. Rn-222 emanation from mine surface areas, and 4. Radionuclides in wastewater discharged to land surface. Estimated average annual dust emissions (item 1 above) from the model mines are listed in Table 2. Emission factors and the assumptions used to estimate these dust emissions are described in detail in the EPA report. Radioactive source terms were computed for each of the sources; dust emissions were multiplied by the concentrations listed in Table 1

Jan 1, 1981

By Dan Nilsson

INTRODUCTION A company must have an objective, and that usually is to earn money. In some countries an objective can also be to reduce unemployment in an area or to develop nonindustrialized areas. Government-owned companies can also have as an objective that of serving the com¬munity. This discussion is limited to the consideration of normal companies wanting to earn money. Such an objective, however, has to be defined in more detail. It is not satisfactory only to maximize the total annual profit from, for example, a mining project. In comparison with most other branches, the mining in¬dustry is very capital intensive. When starting a new project, the mining company must invest a lot of money in mines, ore treatment plants, etc. It is normally many years before any income is received from the ore. Fig. 14 illustrates the payments for a new mining proj¬ect with an expected lifetime of 10 years. Whether this project is a profitable one depends on a lot of things. If the company has a shortage of money, it can be impossible to finance the project. Even if the company has enough money, the management may think that perhaps the project may not return the money fast enough. Such limitations will not be discussed here. Suppose that the company has internal funds from its own operation or can borrow them. Since the mining company must invest the money in advance, it is not satisfactory only to add all of the costs during the mine's lifetime and compare them with the total income during the lifetime. The company must include the time value of the capital. In this appendix, the basic investment theory neces¬sary to judge the profitability and costs of mining proj¬ects will be discussed. INVESTMENT THEORY Time Value of Money If a private person has a dollar and puts it into a bank, he or she will receive interest. If the interest rate is r%, the value after one year will be $(1 + r), after 2 years $(I + r)2 , etc. The interest rate represents the time value of money. If the same person wants to have $1 in the bank af¬ter one year, he or she has to put $(1 + r)-1 in the bank today. If the person wants $1 after 2 years, it is neces¬sary to put $(1 + r)-2 into the bank today, etc. So one dollar today is worth more than one dollar tomorrow. It is, of course, not correct to compare money at differ¬ent times since each dollar has a different value. It is necessary to transform all payments to the same point in time. That is called the calculation of the capital value of a cash flow, and to do this an interest rate is used. In the same way, the cash flow from a project like that shown in Fig. 14 can be translated to any other point in time. Most common is to translate all payments to the beginning of the mining period. That capital value is often called the present capital value. Interest Rate What interest rate will the company use? A com¬pany, just like an individual, has the basic alternatives of keeping its money in the bank and receiving interest, or borrowing money from a bank and paying an interest rate. The interest rate represents the time value of money for the company. If the company has a shortage of money, or many good potential investment projects, it has to use a higher interest rate than if it has a lot of money available. Most companies also increase their interest rate as a hedge against the risk involved by using money for uncertain investments, because of taxes, etc. Capital Value When the interest rate has been decided, it is possi¬ble to estimate the capital value of a project. Here the discussion will be limited to the estimation of present capital values. Suppose that one wants to know the present capital value of $1 (Fig. 15), which will be received after n years. To transform it n years back in time, one must multiply it by the factor (1 + r)-, where r is the inter¬est rate the company uses. Table 1 gives these factors for different years and interest rates. Example: At the end of year 5 a mining company has to invest $50 million in a new mine. What is the present capital value at the end of year 0? The company uses an interest rate of 15%a.

Jan 1, 1982

What is the preferred electrolytic cell design and why? Hall: Considerable research remains to be done on electrolytic cell design. Studies are presently being conducted to determine optimum voltage and amperage. Zadra cells, which were used in some of the first carbon plants, are efficient when properly installed. Gold and silver are deposited on large-capacity, cylindrical steel wool cathodes. At times high value sludge falls from the cathode to the bottom of the cell. Rectangular cells, which were developed at some smaller operations in Nevada, require less floor space and have some desirable features in that cathodes can be moved and sludge can be removed from cell bottoms without actually shutting down the circuit. In South Africa two types of cells using graphite and diaphragm configurations are being tested. Duncan: The rectangular cell has one problem it's easy for shortcircuiting to occur. If you don't have a good seal on the side of the tank and on the bottom, the solution is going to want to flow around the bottom and sides of the cathodes and you're going to get poor performance in the cell. That is the advantage the Zadra cell has over the rectangular cell. With good sealing, which is a must, you should be able to get a ratio of the value of the solution going in to the value of the solution coming out of 10 to 1. Ken, can you tell me what you're attaining on the Zadra? Hall: At the Homestake gold operation in Lead, the cathode in the No. 1 electrowinning cell is loaded to about 600 ounces when it is moved to the refinery. At this time the No. 2 cathode is moved to the No. 1 cell. The No. 3 cathode is moved to the No. 2 cell and a cathode with new steel wool is placed in the No. 3 cell. In rectangular cells this sometimes is not possible because the loading on the cathode in the last cell may be as high as that in the No. 1 cell. As Don has pointed out this can probably be prevented by good seal design between cathodes. Potter: At the risk of sounding as though I'm straddling the fence, I really believe that either the round cell, as developed by Jack Zadra, or the rectangular cell will do an acceptable job. The Zadra is a fine unit. Leaks are not unknown but there are not very many. However, the rectangular cell is very saving of floor space. Plastics have been greatly improved during the past year or two and there is a very efficient design for the rectangular cell. It would have six cathodes in it and contain about 25 ft of steel wool. The body of this rectangular cell would be reinforced fiberglass plastic but with an integral lining of a high temperature plastic. I think either cell is so efficient and reliable that I would almost be tempted to flip a coin depending on floor space requirements. Either cell can be used to pull a loaded cathode and to advance the lesser loaded cathodes. Don, what is the configuration of your cell? Duncan: The Pinson rectangular cell has dimensions 2 ft wide, 30 in deep, and 10 ft long-long enough to take a dozen cathodes and anodes. We are currently using six of each but are not up to full capacity as yet. It's a steel tank coated with rubber. We have copper bus bars, which brings to mind that the last cathode in the tank, in one instance, was loaded heavier than the first one. That was probably a function of poor contact between cathodes and the bus bar. You can have erosion of the bus bar and if you don't get current flowing into the anode and out of the cathode, the cells will not function properly. Kappes: Just a few comments on cell design. In an alcohol stripping circuit we are running in Beatty„ NV, we are using PVC pipes for all circuitry piping and a commercial polypropylene electrolytic cell. They seem to be working quite nicely. The cell itself has eight cathodes measuring 18 in2. At six gallons per minute flow into the cell, at an average gold content of about 500 ppm, we seem to get good loading on the first three to four cathodes and not much on the end cathodes. At 12 gallons a minute flow we seem to be seeing gold uniformly along all eight cathodes. We are running at quite a high amperage level, about 400 amps into that cell. Gene McClelland, US Bureau of Mines, Reno Research Center (from the floor): I was associated with a company where the packing density of their steel wool cathodes was much too high and created some short-circuiting problems in their electrowinning circuit. I would like to ask the panel if they can recommend a packing density for cathodes in pounds of steel wool per cubic foot. Hall: The weight of steel wool put on one of our cathodes is 10 lb and the dimensions of the cathode are such that density would be about 0.75 lb/ft3. Duncan: I made a rough calculation and I'm estimating about 2 lb/ft3. Potter: In checking over a number of electrowinning operations, I found apparently very satisfactory operation within the limits that have been mentioned. I have come to a figure of about 0.5-1.0 lb of medium grade steel wool per cubic foot packed as uniformly as possible, of course. One consideration in packing is the fact that there should be good contact between the steel wool and the conductor. Operators tend to compact the steel wool more than would be desirable just to try to maintain good contact.

Jan 10, 1981

By James R. Arnold, John O. Marsden

Introduction Metallurgical accounting practices have been revolutionized over the past 15 years by the introduction of low cost personal computers and easy- to-use spreadsheet software. The Chimney Creek metallurgical accounting system takes full advantage of these advances. Data is provided by a PLC-based process control system, IBM data logging, operator logsheets and assay results from an automatatic plant sampling system. The data is organized and stored in an IBM PS/2 system using Lotus 123 and customized accounting programs to provide a comprehensive information management system. Metallurgical accounting for gold plants should ideally provide information for the following: a) Gold production accounting b) Day-to-day process control and operation c) Metallurgical evaluation and process optimization The Chimney Creek system provides all of these while overcoming specific accounting problems presented by SAG milling a coarse feed, grinding in cyanide and dump leaching run-of-mine material. Process description Gold Fields' Chimney Creek mine is located 47 miles northeast of Winnemucca, Nevada, at the northern end of the Osgood Mountains. Ore is processed by a combination of dump leaching and milling processes. The dump leach operation treats approximately 2.0 million tons per year of low grade ore (0.030 oz/ton). Run-of-mine ore is top dumped in 40 foot high pads. The use of three solution ponds for leaching enables an intermediate solution to be recirculated to the leach pad to reduce the pregnant solution flowrate and increase gold tenor of solution to the carbon-in-column (CIC) recovery circuit. Gold recovery is approximately 65% by this method. The milling process employs a SAG mill-ball mill grinding circuit to treat 900,000 tons per year at an average grade of 0.19 oz/ton. The ore is crushed to minus 12" in a jaw crusher prior to feeding the SAG mill. Cyanide is added to the grinding circuit to start gold dissolution as early as possible. After grinding, leaching of the ore is continued in a unique two-stage counter current decantation circuit 1 with 12 hours of leaching between stages . 85% of the soluble gold is recovered from the first stage thickener overflow in carbon columns (CIC). The second stage underflow is treated in a scavenger carbon-in-pulp (CIP) plant. Ultimate gold recovery is over 95%. Loaded carbon from both the dump leaching and milling processes is pressure stripped by hot cyanide solution and gold is recovered from the resulting solution by zinc precipitation. The precipitate is retorted for mercury removal and smelted into bullion. The plant flowsheet appears in Figure 1. Data collection Data for daily metallurgical accounting is obtained from three sources: a) Plant sampling b) Process control system c) Operator logsheets All of the data collection for metallurgical accounting is based around the 12-hour shift operating schedule. Sampling All critical sampling for gold production accounting is performed by automatic samplers (Figure 2). Heath & Sherwood two-stage sampling systems are used for sampling all slurry streams (cyclone overflow, CIP feed and CIP tails). The automatic sampler consists of a pneumatically operated primary sample cutter and a continuous rotary secondary splitter. The cyclone overflow and CIP tails are the most critical samples since these are used to calculate gold into and out of the plant on a shiftly basis. The CIP feed sample gives additional information concerning intermediate dissolution and recovery

Jan 1, 1990

BP Minerals America (BPMA) has completed its $400 million modernization of Kennecott Utah Copper's Bingham Canyon mine. Modernized facilities include the in-pit crusher, conveyor, ore storage area, concentrator, ore pipe-line, and tailings pipeline. With modernization, Kennecott will achieve substantial cost reductions. Bingham Canyon is now one of the lowest cost copper producers in the world. And the mine will continue to provide jobs and income in Utah for another 30 years or more. The Modernization Project began in February 1986. The work took 33 months. In late 1987, the construction force peaked at 1100 workers. Kennecott Utah Copper now employs a permanent work force of 2100 employees. Less equipment and more efficiency The modernization work has increased the scale of the facilities that transport and process copper ore. A large, continuous transport system has replaced multiple haulage units. Fewer but larger pieces of equipment have replaced many crushing, grinding, and flotation units. The result is less maintenance, fewer employees, and improved metal recoveries. About one-third of the $400 million modernization project is directed at improving ore movement by eliminating rail transport. The new concentrator accounts for the remaining two-thirds of the modernization investment. The modernized plant was started up in stages. In February, ore was fed to the crusher. By the end of March, the three grinding lines were operating and the flotation circuit was producing copper concentrate. In July, the design through-put rate of 70 kt/d (77,000 stpd) was reached. October saw the achievement of targeted metal recoveries that corresponded to the fully modernized operation. It is estimated that Bingham Canyon will annually produce 181 kt (200,000 st) of refined copper, 9.3 t (300,000 oz) of gold, 71 t (2.3 million oz) of silver, and 5.4 kt (12 million lb) of molybdenite. Current Bingham Canyon production is about 97 kt/d (107,000 stpd) and a similar amount of overburden. Some 70 kt/d (77,000 stpd) of ore are processed through the in-pit crusher and 27 kt/d (30,000 stpd) of ore are transported by rail. The Bingham Canyon mine has yielded more than 11 Mt (12 million st) of refined copper. In 1863, precious metals - silver-bearing lead - were first discovered in Bingham Canyon. What was once a mountain is now a huge open pit. In fact, the mine is the world's largest man made excavation. The pit is some 4 km (2.5 miles) wide and 0.8 km (0.5 miles) deep. Since the pit began operating in 1906, more than 4.5 Gt (5 billion st) of material have been removed. Modernized facilities Electric 21 m3 (27 cu yd) or 23 m3 (30 cu yd) shovels are used to load the haulage trucks. Both 154 t (170 st) and 181 t (200 st) trucks are used to haul the ore to a mobile, in-pit crusher. The crusher replaces three crushers located at the old concentrators. There are two dumping locations at the new crusher, which can handle two trucks a minute. The 1.5 x 2.7 m (60 x 109 in.), 745 kW (1000 hp) Allis-Chalmers gyratory crusher reduces the ore to 254 mm (10 in.) size or less and delivers it to the conveyor. The crusher and conveyor systems are capable of handling 9 kt/h (10,000 stph) of ore.

Jan 11, 1988

By T. Gupta, T. Ghosh

"In recent years, the demand for Rare Earth Elements (REE) have grown rapidly due to increasing demand and global supply shortage. The supply deficit of these critical elements has encouraged the search for new sources of REEs, possibly Coal and Coal byproducts. Coals from certain parts of the world can be rich in REEs, and can approach a total concentration of 1000 ppm. Two Alaskan coal samples were investigated for the effect of size and density on the concentration of REEs on three selected size fractions. Flotation tests were conducted for finer fractions. Additionally Bottom ash and Fly Ash samples from a power plant have been examined for REE concentration. The results show that the upgrade potential for REEs on ash basis from whole coal basis ranges from 2:1 to 4:1 for Wishbone Hill and Healy coal samples respectively. Flotation of the finer fraction for both coal samples, conducted under similar conditions, reveal higher concentration of REE’s in the tailings. Both coal samples have a comparatively higher LREE content than HREE. Trends of REE content of the power plant products on ash basis revealed that Fly Ash has a slightly higher concentration of both LREE and HREE than Bottom Ash. Keywords: Rare Earth Elements, Coal, Coal Ash INTRODUCTION Rare Earth Elements (REEs) and their compounds are vital components in industrial, automotive, petrochemical, defense and medical applications. Dysprosium in combination with other elements is used in laser materials and nuclear control applications. Neodymium is present in magnets that are used in green power generation such as the manufacture of wind turbines and also in hybrid cars. Europium, Terbium and Yttrium are used for devices that have visual displays such as computer screens and laptops [1, 2]. REEs pose fewer environmental problems of disposal and recycling than conventional heavy metal counterparts [1]. Rare-earth metals are key ingredients for radar systems, avionics, modern national defense and satellites [2-8]. In the last 25 years China has been mining, processing and refining Rare Earth elements at affordable prices thus gaining a virtual monopoly in the market. The many fold rise of the prices of REE elements over the last few years and frequent disruptions in supply has forced various nations to invest in indigenous extraction of REEs by increasing production from previously productive sites and to expand the production from new alternate sources including coal and coal byproducts [9]. Coal and Coal beds have been found to be significantly enriched with certain critical metals and can be a potential source for both Light Rare Earth Elements (LREE) and Heavy Rare Earth Elements (HREE) [10]. Moreover, coals with high REE concentrations have been discovered in Pavlovka and Rakovka coal deposits in Russia, Appalachian deposits in USA, Sichuan Basin deposits in China and others across the globe [10-13]. A recent study from China looked at how different trace elements (including numerous REEs) partition in various size fractions of Antaibao coal from China. The finding indicates that the REEs, phosphorous, and thorium all have their high concentrations in the particle sizes 0.5 to 3 mm, and 6 to 25 mm. REEs are also often present in the form of phosphates such as monazite in coal [13, 14]. The objective of the study is to identify the concentration of REEs in Alaskan coals and to investigate the effect of density and size in the distribution of REEs in coal and coal byproducts."

Jan 1, 2016

By Syd S. Peng

According to accident statistics (l, 2), roof fall accidents occur quite frequently at the intersections of underground? openings and account for up to 30% of the roof fall fatalities. But this figure will certainly go up considerably if roof falls not associated with any fatal accident are included in the analysis. Unfortunately, nonfatal roof falls are seldom documented. In the underground coal mines, there are two types of entry intersections. One is the three-way (V type) and the other is the four-way (T type) intersections. A three-way intersection occurs whenever the mainline track haulageway turns at an angle into a section haulageway. The four-way intersections by far the most common type in the U.S. under-ground coal mines are formed by the intersections of two entries which are oriented perpendicular to each other or by the intersections of entries and crosscuts. However, crosscuts are frequently driven at an angle to the entry to facilitate machine movements. The four-way intersections formed by this method of crosscut development are angled. At the intersection of two underground openings, the diagonal roof span is wider than the width of each individual opening. This is one of the major reasons why roof falls occur more likely at the intersections than in the entry between pillars. Consequently, special intersection support plans are usually employed (Figs. 1 and 2) for intersections. The most popular practices are to decrease roof bolt spacing 1 ft. less than or to increase roof bolt length 1 to 2 ft. longer than those employed in the entry between pillars (3). The increase in bolt length or decrease in bolt spacing can be applied to all of the bolts in the intersection or employed only for the strategically located bolts depending on the individual mine practices. These rather generalized practices seems to be adequate for some areas but in many cases inadequate for others. Occasionally, the neighboring two lines of chain pillars are staggered such that the crosscuts in one row of chain pillars face the centers of the chain pillars in the neigh-boring rows. This layout forms T-type three-way intersections which are also inadequate for most weak roofs (3, 4). A study was therefore initiated to develop special techniques for intersection supports. The first phase of the study was the analysis of the stress fields for the underground entry intersections. A three-dimensional stress analysis was necessary for this purpose. Unfortunately, the three-dimensional problems of complicated boundary conditions are beyond the existing analytical methods. The three-dimensional finite element method (FEM) was therefore employed in this study. In the analysis, square pillars of 40 feet wide and an entry height of 8 feet were assumed while entry widths varied at 14, 20 and 26 feet. This model was selected to simulate the mine where roof falls data were collected. The second phase of the study was the collection of field data. The sizes of the roof falls including heights, lengths and widths were collected and analyzed. In the third phase study, the results of the first and second phases were compared and discussed. The analysis of field data revealed an arching zone within which roof falls occurred. The arching zone above an entry obtained empirically was compared with the calculated stress contours. Assuming that the bolts must be anchored in the rock strata outside the arching zone, the preferable systems of bolting were proposed. Also the amount of pretension required to inhibit sliding between layers was discussed.

Jan 1, 1978

By M. R. Yalamanchili, C. A. Young, J. J. Kellar, W. M. Cross, J. D. Miller

Introduction Researchers have long speculated regarding the adsorption state of collectors on mineral surfaces. Only recently have analytical techniques become available that allow direct in situ analysis of adsorbed layers. With reactive internal reflection elements (IREs), Fourier transform infrared spectroscopy coupled with internal reflection spectroscopy (FT-IR/IRS) allows for unobtrusive investigation of the adsorbed layer (Harrick, 1979). This technique has been used both for the in situ identification of adsorbed species and for the direct calculation of collector adsorption density in real time from infrared spectral data (Miller and Kellar. 1988). The reactive mineral IRE serves both to internally reflect the infrared light and to act as a substrate at which collector adsorption can be monitored. Shown in Fig. I is a schematic of light being internally reflected in an IRE. As the light is totally reflected at the interface, an evanescent wave propagates a short distance (typically a few microns) away from the IRE, and sampling of the interfacial region occurs. Because of multiple internal reflections, great sensitivity can be obtained. Flotation systems for which this method has been demonstrated include CaF2/oleate, Al2O3/dodecylsulfate, CaCO3/oleate, ZnS/xanthate and KCI/octylamine (Miller and Kellar, 1988; Kellar et al., 1989, 1990; Miller et al., 1990). Of particular interest to flotation chemists is the conformational behavior of the alkyl chain of adsorbed collector molecules. In the simplest sense, conformation can bethought of as the relative flexibility of the hydrocarbon chain caused by rotations about carbon-carbon single bonds. The flexibility is described by the amount of gauche/trans conformations in the alkyl chain. Figure 2 shows examples of gauche and trans conformations (rotational isomers or rotamers) in a hydrocarbon chain. These two conformations are observed when the terminal methylene groups in any four-carbon subchain are on the same side (gauche) or opposites sides (trans) of the plane of the middle carbon-carbon bond. A gauche conformer (rotamer) is thus a state of maximum potential energy, whereas a trans conformer is a state with minimum potential energy. The difference between these states has been found to be approximately 0.5 to 0.8 kcal/mole (Mead et al., 1986). The relative number of butane molecules present as gauche conformers with respect to trans conformers can be calculated from the Boltzmann equation (Mead et al., 1986). At 25° C approximately 66% of the butane molecules are trans rotamers. [ ] However at 100° C, only 33% of butane molecules are trans rotamers. Exposing the hydrocarbon chain to a polar, rather than an a polar, medium should also cause an increase in the number of gauche butylene groups in a hydrocarbon chain, although this effect has been difficult to observe experimentally (Mead et al., 1986). Forming a crystalline hydrocarbon is equivalent to decreasing the temperature, which greatly increases the amount of trans butylene groups present in the chain. From the preceding discussion, a picture of the solution species of concern can be developed. The collector monomer should exhibit the greatest number of gauche rotamers and most likely will look crumpled or balled-up to reduce the surface area in contact with water. The micellar form is expected to be aggregated, with the hydrocarbon chain existing in a less polar environment. These characteristic conformational features have been examined by FT-IR spectroscopy for various liquid phases of alkanoic acids (Umemura et al., 1981; Cameron et al., 1982a; Yang et al., 1986; Wong, 1987; Wong and Mantsch, 1983; Cross et al., 1992), phospholipid membranes (Fringe Ii, 1980; Casal and Mantsch, 1985: Wong et al.. 1988;

Jan 1, 1994

By J. D. Morgan

Demand for nonfuel minerals depends in large measure on economic activity, which rose in each quarter of 1985. However, despite 1.7 million new housing starts and a 7% increase in motor vehicle production, many domestic metal mining operations experienced continued difficulties as a result of import competition, greater use of newer engineered materials, emphasis on multi-unit housing, and downsizing of motor vehicles. Recognizing the continued difficulties of the domestic metal mining industry, many corporations made major efforts to retrench and rationalize - selling, closing, or spinning off less profitable operations, renegotiating labor contracts, and inviting increased participation by foreign firms. About 15% of US energy was imported in 1985. This included $52 billion worth of crude oil and refined products, adding significantly to the overall merchandise trade deficit of $149 billion. With the exception of a few large-bulk, low-value raw materials, most minerals move relatively freely in world trade. The strong dollar encouraged imports and made exporting more difficult US nonfuel minerals trade deficit totalled $17 billion. The US continued to rely on imports for many important strategic minerals in 1985 (see accompanying Table). Increased pressures for protection included major mineral products such as steel, ferroalloys, and copper. But President Reagan, in his Feb. 1986 "Economic Report to the Congress" stated: "Our international trade policy rests firmly on the foundation of free and open markets." He also stressed the need for fair trade and more open markets internationally. The National Critical Materials Council, created by law in the latter part of 1984, was activated in the latter part of 1985. Its first public hearing was held in January 1986. The Council is to review the nation's materials posture and advise the President on such matters. Compared with 1984, the value of US raw nonfuel mineral output rose 2%, to $24 billion. This supported the production of $244 billion worth of processed materials of mineral origin. Raw steel production - a major indicator of both material use and mineral production - fell 4% to 81 Mt (89 million st). Consequently, domestic iron ore production fell to 49 Mt (55 million st), about 97% pelletized. Iron ore imports fell 7% to 16 Mt (18 million st). Raw steelmaking capacity fell 4% to 117 Mt (129 million st). Pacts with 16 leading steel exporting nations have been negotiated in an effort to reduce import penetration of the domestic market. Domestic producers of bulk ferroalloys of chromium, manganese, and silicon operated at about 40% capacity. The US mined no manganese or chromium. Only a small quantity of ferronickel was produced from domestic laterite mining in Oregon. Molybdenum production increased slightly to 49 kt (54,000 st). Mine shipments of tungsten, another mineral tied to steel production, also fell slightly to 1.1 kt (1200 st) of contained metal. Even the domestic aluminum smelting industry suffered reverses as high stocks and low prices reduced production 16% to 3.5 Mt (3.8 million st). Similarly, imports of bauxite fell 7% and alumina 10%. Production of magnesium metal fell 6% to 136 kt (150,000 st), while titanium metal production remained unchanged at 22 kt (24,000 st). Copper mine production was unchanged at 1.1 Mt (1.2 million st). Copper refinery production fell slightly to 1.1 Mt (1.2 million st). Lead mine production rose 24% to 400 kt (440,000 st). Zinc mine output was down 10%, to 225 kt (250,000 st). Refined lead production increased 28%, to 510 kt (560,000 st). Primary slab zinc production fell 5% to 240 kt (260,000 st).

Jan 5, 1986

By Jozef S. Stachulak

INTRODUCTION Conveyor belt fires represent a serious hazard in underground mines. This fact is well understood in most mining countries. The statistical data for Ontario mines (1) indicates that the majority of the conveyor fires occurred at the head pulley of the conveyor belt. The main emphasis for fire prevention and detection should then be concentrated at the drive pulley of the conveyor designed to transport non-combustible material. In the case of conveyor belt fires, it is of utmost importance to detect the hazard at an early stage, when the overheating has not yet reached the ignition point. Thermal noise detection sensors were installed in 1987 at several conveyor sites at Inco mines. PRODUCTS OF COMBUSTION THERMAL DEGRADATION OF CONVEYOR BELT Various investigations on conveyor belt fire hazards have been made in the past (2, 3, 4, 5, 6, 7). Several belt samples (6, 7) were subjected to thermal oxidative degradation. The belts used in these tests were neoprene, polyvinyl chloride and styrene-butadiene rubber (6). About 70 different compounds were identified in the volatile thermooxidative products. As expected, hydrogen chloride was the major toxic product formed, accompanied by relatively low quantities of other toxic compounds such as carbon dioxide, carbon monoxide, sulfur dioxide, carbon disulfide, hydrogen sulfide, benzene, chloroprene, cresols, chloroethanol, formic acid and C8 hydrocarbons. The research investigation (8) performed by the USBM indicated that "smoldering combustibles generate a very low level of CO compared to the level of smoke produced". The cited publication (8), states that, "During smoldering combustion, typical mine combustibles produce about 2.5 times more smoke than CO. As fires develop to their flaming stages, the trend reverses and about eight times as much CO is produced as smoke." The reader is referred to several significant sources of additional information pertaining to emission products that can be found in references (8), (9), and (10). ANALYSIS OF FIRE HAZARDS U.S.A. AND ENGLAND Nearly 15 percent of all underground coal mine fires in the United States involved ignition of conveyor belts (7). In many cases, the fires propagated through thousands of feet of belt entry and created losses in the millions of dollars. Some of these fires consisted of rubber belts. Some involved polyvinyl chloride and neoprene belts that were classified "fire resistant". Typically, a belt fire begins after the conveyor is stopped. The fire is initiated by burning coal dust and grease accumulation around defective rollers or belt drives that have become heated by friction (7). An analysis of 113 conveyor fires (11) showed that 41 originated at the driving pulleys, 32 at the defective rollers, and 40 elsewhere on the conveyor system. The severity of the fire, which can spread along a conveyor road, has been emphasized by the disaster at Creswell Colliery, Derbyshire, on September 26, 1950, when 80 men were overcome and killed by toxic fumes from the fire (12). The fire started from smoldering caused by friction, and developed into a flaming combustion, propagating along the roadway containing the rubber belt at an estimated speed of 40.2 m/s (90 mph), and travelling a total of 650 m (2100 ft) THE INFLAMMATION OF CONVEYOR BELTS THE MECHANISM OF IGNITION Defective operation of a conveyor belt may start a fire in several ways: - Friction of the belt on the drive pulley. - Jamming of a guide roller, and others.

Jan 1, 1991

By P. T. Bissert, T. J. Lutz, J. A. Yonkey, G. T. Homce

"The U.S. National Institute for Occupational Safety and Health (NIOSH) has been researching refuge alternatives (RAs) since 2007. RAs typically have built-in pressure relief valves (PRVs) to prevent the unit from reaching unsafe pressures. The U.S. Mine Safety and Health Administration requires that these valves vent the chamber at a maximum pressure of 1.25 kPa (0.18 psi, 5.0 in. H2O), or as specified by the manufacturer, above mine atmospheric pressure in the RA. To facilitate PRV testing, an instrumented benchtop test fixture was developed using an off-the-shelf centrifugal blower and ductwork. Relief pressures and flow characteristics were measured for three units: (1) a modified polyvinyl chloride check valve, (2) an off-the-shelf brass/cast-iron butterfly check valve and (3) a commercially available valve that was designed specifically for one manufacturer’s steel prefabricated RAs and had been adapted for use in one mine operator’s built-in-place RA. PRVs used in tent-style RAs were not investigated. The units were tested with different modifications and configurations in order to check compliance with Title 30 Code of Federal Regulations, or 30 CFR, regulations. The commercially available relief valve did not meet the 30 CFR relief pressure specification but may meet the manufacturer’s specification. Alternative valve designs were modified to meet the 30 CFR relief pressure specification, but all valve designs will need further design research to examine survivability in the event of a 103 kPa (15.0 psi) impulse overpressure during a disaster. BackgroundThe Mine Improvement and New Emergency Response Act of 2006, known as the Miner Act (U.S. Department of Labor, 2006), was enacted in the wake of three mine accidents involving explosions or fire that claimed 19 lives that year. Intended to help improve underground coal mine accident preparedness, the Miner Act includes provisions that target mine safety issues in areas such as emergency response planning; adoption of new technology; training and education; and mine safety standards enforcement. Section 13 of the Miner Act specifically directed the U.S. National Institute for Occupational Safety and Health (NIOSH) to provide for research into the effectiveness and viability of refuge alternatives (RAs) for underground coal mines, and the U.S. Department of Labor to act on the results of such research, as appropriate. These mandates culminated in the 2009 adoption of changes to the Title 30 Code of Federal Regulations, or 30 CFR, mining health and safety regulations (U.S. Department of Labor, 2008), requiring underground coal mines to supply mine emergency RAs and associated components so as to provide a life-sustaining environment for persons trapped underground. Such RAs can be either selfcontained mobile units or built-in-place facilities. The regulatory changes also include provisions establishing requirements for the approval of RAs and their components by the U.S. Mine Safety and Health Administration (MSHA), among which are numerous criteria for providing a safe breathable atmosphere under positive pressure within the RAs. One criterion for maintaining a safe RA atmosphere is the inclusion of an air pressure relief valve that will activate at a maximum of 1.25 kPa (specified as 0.18 psi, or approximately 5.0 in. H2O), or as specified by the RA manufacturer, above mine atmospheric pressure in the RA (U.S. Department of Labor, 2008)."

Oct 1, 2016

By Stanislaw Siewierski

One of the major problems in the exploitation of deep copper ore deposits in the Legnicko-Glogowski Copper Field is ground control and, in particular, roof control of the headings. The depth of copper deposits ranges from 600 to 1200 m. The parameters of ore-bearing rocks and the surrounding rocks show considerable variance of strength properties along the mining thickness (Fig. 1). The roof of the openings exploitation headings consists of dolomite-calciferous rocks. The roof strata consist of layers, the thickness of which vary from several centimeters to 0.6 m (sometimes to 1.0 m). These strata differ in strength and are characterized by igneous and sedimentary discontinuities. On the basis of studies in structure, properties and geological mining conditions of the roof strata in Polish copper mines, the method of roof classification has been developed. (1) This method uses the stability index "L" of the roof, calculated from [ ] The value of "t" ranges from 0, l to 0,8. Based on the value of "L", four classes of the roof (Table 1) corresponding to different degrees of fall hazard have been developed. The mining of copper ore deposit is carried out by a single lift room-and-pillar mining system or a two-stage room-and-pillar system with roof caving. (3, 4). The single lift room-and-pillar mining system is most often used. It consists in working of the deposit of 2.5 to 4.5 m mining thickness, 150-350 m long mining line, consisting of parallel room faces 6 m wide with pillars 5.7 or 16 m wide. As the mining line advances, cross-cuts joining individual rooms are made, thus forming support pillars 3x5, 7x7, 5x16, and 7x16 m (Fig. 2). The two-stage system (Fig. 3) mines the deposit with rooms 6 m wide into blocks 25 x 35 x 25 x 45 m. During the second step these pillars are worked with rooms and cross-cuts into pillars about 3 x 4 to 4 x 5 m (retreat mining). The pillars are crushed just before the fall of the roof. The maximum opening of workings in the two-step system is about 10 m, i.e. one belt entry and one row of bearing pillars. In single lift room-and-pillar systems it is about 35 m., i.e. three belt entries and three rows of bearing pillars. The support of the headings consists of expansion bolts 1.6-1.8 m installed in a bolting pattern 1.1 x 1.5 m. In both systems the fall step is about 10 m, i.e. the width of a belt and a bearing pillar. The Indications of Rock Mass Pressure Individual Dog Heading Individual dog headings are set wide apart from each other in the unmined ore deposit area where there is no interaction of headings. The indication of rock mass pressure are, in such cases, limited to the deformations and damage of side walls. Radiometric studies of rock mass loosening showed that the maximum range of lifts in side-walls immediately after web was 2 m (4,5). In the course of time the cracked side rocks fall off, thus widening the heading. The morphology of side-walls is influenced by the location of the heading in relation to clump layer/tarry. This layer is characterized by a weak strength and "lubricating" properties. The location of the heading in

Jan 1, 1982

By John R. Alderman

INTRODUCTION Uranium production in Saskatchewan commenced in 1953 at the Eldorado Nuclear Mine in the Beaverlodge area of North Saskatchewan, followed over the next few years by nineteen other smaller mining operations. Production rose from 453 600 to 2 721 600 kg per annum by 1958. Over supply and the closure of the U.S. market to Canadian uranium led to an equally rapid decline to an annual production which varied from 453 600 to 907 200 kg during the years 1964 to 1976. As the market strengthened there was a resurgence of the Saskatchewan industry and production from one underground and two open pit mines reached 2 809 000 kg in 1980 for a return of 232 000 000 dollars (Table 1). [ ] Future output will be increased by a potential 4 536 000 kg from the Key Lake mine now under construction and scheduled to commence production in 1983. Three other projects are in an advanced stage of planning and are currently in the regulatory review process. Several other prospects are being evaluated. Figure 1 shows the location of uranium mines in Saskatchewan and table 2 illustrates the present status of uranium mines and deposits in Saskatchewan. At present, 1400 workers are employed directly by the uranium mining industry. Saskatchewan has diverse mining conditions ranging from the underground mine of Eldorado Nuclear Limited 1870 m deep with a low grade ore deposit averaging from 0.17-0.20% with extraction by carbonate leaching through an open pit operated by Gulf Minerals with a moderate depth of 110 m and 0.37-0.5% grade to the shallow open pit of Cluff Mining 30 m deep with a very high average ore grade of 7.0%. Extraction at both open pits utilizes acid leach extraction processes, the latter after gravimetric separation. Key Lake mining has commenced development of the next mine to come into production which will have two open pits to a depth of 80 and 175 m with an average 3% grade to be followed by the deep open pit of Canada Wide Mines Ltd. to a depth of 205 m with average grade of 1.5%. One existing mine will transfer its operation from the present open pit to another of slightly higher average grade. From the Saskatchewan point of view, these different operating conditions present varied problems associated with occupational health and safety and concomitant solutions to reducing worker exposure to radiation. REGULATION OF URANIUM MINES In 1946, the Government of Canada enacted the Atomic Energy Control Act which established the Atomic Energy Control Board (AECB) and vested in that board powers to regulate all aspects of uranium mining. In the early years, the AECB relied heavily on provincial occupational health and safety agencies for the surveillance of uranium mining and milling operations, particularly with regard to "conventional" occupational health and safety, the AECB having at the time no specific regulations covering the mining industry. There has been considerable dialogue between federal and provincial governments since, traditionally, occupational health and safety is a provincial jurisdiction. Saskatchewan is strongly opposed to the intrusion of the federal government into the provincial domain, particularly as there is little justification for a federal presence in occupational health and safety in uranium mining and milling. The province does, however, accept the need for federal involvement in certain aspects of the regulation of nuclear industry, for example, international sales, research and security of reserves for domestic use.

Jan 1, 1981

By Prescott Greene

AVAILABLE MACHINES Machines discussed in this section include only rubber tire mounted kerf cutting machines of a self¬propelled electrically powered design. Such machines handle bars up to 4.6 m (15 ft) in length, cutting kerfs of from 115 to 191 mm (4.5 to 7.5 in.). The basic design concept of all machines currently in production is very similar. The chassis features dual driving wheels on the cutter-bar end, fixed directly to the main frame. The steering wheels on the outby end of the frame are articulated in some manner to keep all four wheels on the floor over uneven bottom. Machine power for tramming, and all other functions except powering the cutting chain, is furnished from a pump motor that drives the specific pump or pumps required by a given machine. The cutter chain is driven through a mechanical drive train from a separate electric motor having no duty other than driving the chain. The general arrangement of a cutting machine is shown in Fig. 1; two types of machines, with their specifications are shown in Figs. 2 through 5. A hydraulically powered cable normally is included to eliminate the need for manual cable handling and make this a one-man machine. A bug duster, though optional, is needed for most applications. The bug duster is an hydraulically driven scroll conveyor located directly behind the cutter chain drive sprocket. Its function is to move the cuttings away from the sprocket to prevent them from packing up in this area and consuming a lot of machine power to further pulverize them. Available machines vary from models 610 mm (24 in.) high that can cut horizontally only from 203 mm (8 in.) below ground line to 229 mm (9 in.) above, to machines 1143 mm (45 in.) high that can cut in any plane from 178 mm (7 in.) below the ground line to 3.7 m (12 ft) above it. Most currently manufactured machines fall in the 17 200 to 20 400 kg (38,000 to 45,000 lb) weight range, but can be dis¬assembled for underground installation where weight and size are a problem. Maximum teardown results in all pieces having a weight of 5400 kg (12,000 lb) or less and a size allowing passage through a 1.2 x 2.1-m (4x7-ft) opening. Further reduction in either size or weight of the parts becomes a serious problem. Machines currently being manufactured are avail¬able for use with power supplies of 250 v dc; 500 v dc; 440, 550, or 950 v ac (60 Hz). Also available are machines for use on 415, 500, and 950 v 50 Hz. De¬signs to allow operation on other voltages of either 50 or 60 Hz can be made available without any serious problems. As of September 1977, a typical medium-seam ma¬chine without the extra options needed for unusual conditions, but fully equipped to operate, was selling for about $185,000. CAPABILITIES Materials That Can Be Cut The problem of deciding before a machine is put on the job whether it can work effectively is an old one that has never been completely solved and probably

Jan 1, 1982

By Deepak Malhotra, Linda Battich

INTRODUCTION Mineral processing plants generally operate efficiently within the technical, economic and management constraints applied to them. They have no choice, because the dynamic and competitive nature of the mineral business dictates that survival and/or prosperity of the company depends on being a low cost producer. Progressive mineral companies realize that market downturns will occur, and they must continuously strive for lower unit production costs rather than be satisfied with the status quo. Mineral companies generally tend to reduce costs during market downturns. This short-term solution is achieved by high grading of ores, and cut-backs In waste stripping, exploration, and capital expenditures.(1) However, to remain a low-cost producer in the long term, companies must enhance revenue and reduce cost simultaneously. This Involves continuous evaluation of technologies and reagents, periodic audits to locate revenue losses; and gathering and analyzing statistical plant data to clearly understand the Impact of process variables on efficiency. Mineral companies are caught in a Catch-22. They need resources to accomplish the objectives outlined above. Most mill stuffs have been cut to the minimum level necessary to maintain production. Resources such as time, manpower and money are scarce, and mill metallurglsts struggle to use these resources productively. Is your flotation plant optimized? History Indicates that successful operations constantly tune the flotation circuits to Improve metallurgy or reduce cost. This Is the essence of evolutionary change In mineral processing operations. A plant auditing approach can play a key role in assisting the process engineers to achieve the overall objective of enhancing metallurgy and reducing cost with limited resources through evolutionary improvements. FLOTATION SYSTEM The flotation process is the most widely wed concentration process In the world today. The number of variables that can affect the flotation results Is very large, maybe over 100. These variables can be divided Into three main groups: (2) (1) ore and mineral properties; (2) reagent treatment; and (3) flotation machine characteristics. Some of these variables can be controlled while others have to be accepted. Recently, Investigators (3,4) have attempted to describe the flotation process as an engineering system which includes chemistry, equipment and operating variables. Though some of the variables are Independent, most are co-dependent. thus making it difficult to develop simple cause and effect relationships. For example, it is easy to understand the effect of reagent change on recovery and grade In the rougher circuit, but very difficult to quantify Its impact on filtration downstream. Plant audits on a periodic basis are necessary to develop an understanding of the Interactive nature of the system. Auditing provides comparisons of plant operating performance before and after variable changes. OVERVIEW OF PLANT AUDITING An audit Is a formal, thorough and periodic examination of a system. There are various types of audits (i.e., geological, environmental, metallurgical. economic, etc.), and each serves a different purpose. (5) The discussion here is related to plant and/or metallurgical audits. Neale and Flintoff note that two key factors governing changes In an operation are: (1) the opportunity must be recognized; and (2) the justification for change must exist. (6) Both factors emphasize the critical role of successful plant auditing. Without plant audits. many opportunities for improvement must go unrecognized. On the other hand, when a situation ripe for improvement Is recognized. the hard data from a plant audit can reveal the expected impact of an operating change.

Jan 1, 1993

By B. J. Arnold

Froth removal rate is perhaps the most important parameter in scaling up from laboratory tests to plant froth flotation circuits. The WEMCO froth factor, the percentage of the froth height over the lip of the cell, accounts for the difference between laboratory and plant froth removal rates. As laboratory tests are generally run with a high pulp level, laboratory results are directly comparable to plant results only when a high froth factor (high pulp level) is run in the plant. When a plant does not have a high pulp level, laboratory results must be modified by using a ratio between laboratory and plant kinetics to predict plant results. This ratio, which accounts for the froth factor is important in testing any new reagents or reagent dosage levels. CQ Inc. researchers have used predictive scale-up equations using this froth factor coupled with equations for kinetics and the residence time distribution in the froth flotation cells. These equations show that modifying the froth factor in the plant can have a large impact on flotation cell yield and reagent use. EQUATIONS DESCRIBING FROTH FLOTATION Scaling up from laboratory- to plant-scale coal froth flotation has been the subject of much research and debate over the years. Determining the proper rate equation to describe flotation kinetics and an equation to describe the residence time distribution in a bank of flotation cells have been particular areas of study. As shown in Tables 1 and 2, various rate equations and residence time distribution equations have been proposed in the literature. These equations consist of one or two parameters giving the rate constant(s) in units of time-1 and a parameter for recovery at infinite time, [Roo]. Dowling, Klimpel, and Aplan (1985) evaluated the rate equations in Table 1 for their ability to fit froth flotation kinetic data and found that all of the equations give reasonably low standard errors. Several of these equations, however, proved to have superior confidence limits for the parameters. Other equations have also been proposed to describe the recovery versus time profile. Most recently, Diao et al. (1992) proposed a first-order form with a sine wave distribution. As indicated in Table 2, investigators have suggested many equations to describe the residence time distribution in continuous froth flotation. To measure the water residence time distribution in a bank of cells, a soluble tracer is often used, with the level of tracer being measured over time at the tailings discharge. This response is then put on a dimensionless basis as described by Austin et al. (1984) in order to compare mixing behavior between cells or under different conditions. Coal flotation cells are generally of the flow-through or hog-trough variety allowing for some degree of backmixing and short-circuiting. While the water residence time distribution (see Figure 1) at the end of one cell in a bank of 1.4 cu ft WEMCO froth flotation cells may approximate perfect mixing conditions (Arnold, 1989), the water residence time distribution at the end of a bank of cells does not. Figure 2 shows the RTD for various flotation cell types compared to that for perfect mixers in series. To better describe the water residence time distribution in coal froth flotation cells, Van Orden (1986) used an equation similar to the perfect-mixers-in-aeries equation. This

Jan 1, 1993