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By M. K. Quinn, A. W. Khair

"This paper presents an analysis of respiraple dust generation due to the action of a continuous miner. Underground coal cutting by a drum-type continuous miner was simulated in the laboratory using a specially designed unique automated rotary coal cutting simulator (ARCCS). An analysis of -400 mesh particles was done on gravity collected material. Samples of entrained dust were collected through the use of cascade impactors. Both samples were analyzed for size distribution and particle shape. Using the simulated coal cutter permitted variation of operating and in-situ parameters to provide information on the effects on generation of -400 mesh particles. In these tests coal blocks with an approximate dimension of 18 in. x 15 in. x 6 in. (45.7 cm x 38.1 cm x 15.2 cm) were first subjected to confining pressures, equivalent to in-situ conditions. Such blocks were then cut by the cutting head of the ARCCS, thus simulating the action of the continuous miner in underground coal mining. During the tests under a particular set of in-situ and operating parameters a number of other parameters such as 1) penetration of bit into coal, 2) penetration resistance (thrust and cutting pressures), 3) rotating velocity of cutting head, and 4) acoustic emission activity in the coal block were monitored. After each cutting cycle the fractured surfaces were photo¬graphed and a velocity survey was conducted by using a sonic technique. At the end of each experiment the cutting paths of the bits in coal were photographed using an optical microscope with an attached camera.IntroductionThe continuous-mining machines introduced in the 1950's now account for more than half the production of coal from underground mines. Unfortunately, these continuous miners, designed for increased productivity, have also increased the concentration of respirable dust in mines. As a result, the Federal Coal sine Health and Safety Act of 1969 was enacted to enforce on the coal operators that the airborne respirable dust not exceed 2 mg/m3, and was intended to reduce the incidence of coal workers pneumo-coniosis. This plus the other aspects of the legislation and the subsequent monitoring and enforcement by MSHA made it very imperative that the coal operators provide a relatively healthier and safer working environment for the coal miners. Since 1970, the Federal government has paid over 11.7 billion dollars to more than 470,000 miners with coal workers pneumoconiosis and their survivors (Newmeyer, 1981). These regulations, coupled with the assessment of the continuing burden on miners, the mining industry and taxpayers appear to have provided an impetus for the mining community towards comprehending the different parameters that influence dust generation and entrainment."

Jun 1, 1988

By K. G. Wheeland, K. G. Byberg

The occurrence of a secondary, dust-fuelled explosion following blasting of sulphide ores has been reported in mines in many countries, including canada, the USA, Australia and South Africa. However there has been little effort to document the events and develop effective control technology. Following a detailed investigation of an incident at the Geco Di vision of Noranda Minerals Inc. , the company held a conference of interested parties in October 1986 to review the state-of-the-art and to agree on action. '!his paper outlines the causes and effects of such explosions, the key conference points and conclusions, and subsequent information-exchange and teclmology-developrnent activities.

May 11, 1988

By A. Bauer, P. Katsabanis, M. Baker

The principal results from a previous paper presented at the Mine Accident Prevention Association of Ontario in 1987 (1) are summarized here. It is concluded that the selection of explosives in mines experiencing sulphide ignitions is important and that aluminized explosives should be avoided because of their high explosion temperatures and their detonation products. The duration of the blast is also critical and secondary blasts should not occur simultaneously with nearby priroa:ty blasts.

May 11, 1988

By G. A. Moruzi

Two sulphide dust secondacy explosions have occurred in the mines of the SUdbury operations of Falconbridge Ltd. during the past 18 months. 'Ihe main similarity between the occurrences is that they both happened in vertical crater retreat blasthole stopes. A relatively high p:,v.tier factor was noted in both instances. other details are enumerated and examined, as well as some results of tests done by Noranda Research on company ore samples.

May 11, 1988

By Q. Liu, P. Katsabanis, J. A. Meech

A thermohydrodynamic computer code has been used to simulate dust explosions. For sulphide dusts, the thermodynamic predictions are used to investigate the effect of dust concentration, mineral content and inert materials as well as the quantity of sulphur dioxide produced. It was found that the amount of oxygen available for the explosion is very important and that the dust concentration which gives maximum explosion pressures is obtained at zero or slightly negative oxygen balance values. The sulphide minerals were found to have the following explosibility sequence: pyrrhotite, pyrite, chalcopyrite, sphalerite, covellite, chalcocite and galena, which is in agreement with mine experience. The results show that theoretical calculations can be used to predict explosion pressures and toxic gases produced by a dust explosion and therefore they can assist in the design of protective measures .

May 11, 1988

By M. J. Sapko, R. W. Watson, E. S. Weiss

The relative incerdivity of several blasting agents was evaluated in the Bureau's Bruceton Experimental Mine (BEM) . The minimum charge weights resulting in the ignition of predispersed oil shale dust clouds were measured for each of the explosives. A 400 g unstennned charge of ANFO (ammonium nitrate-fuel oil) was sufficient to initiate an oil shale dust explosion, whereas a 1000 g unstenuned charge of an emulsion blend did not ignite the dust even with the addition of 4% methane gas. Several stenuning materials were evaluated in the BEM for their effectiveness in preventing the ignition of near stoichiometric methane-air mixtures. Water stenuning was superior to all other inhibitors tested against methane. Explosibility testing of fine sulfide ore dust in a 20 L chamber and in the BEM show that the dust is capable of propagating an explosion given nominal concentrations of 700 to 800 g/m3. Results are encouraging and irrlicate that there is merit in pursuing explosive composition as a means of reducing ignition probability of oil shale and sulfide dust clouds generated during nonnal blasting operations. The proper selection of explosives and stenuning materials should significantly reduce the probability of secondacy dust explosions in oil shale and sulfide ore mining.

May 11, 1988

By R. J. Enright

Eight mines in canada and three in Australia were visited between September 1987 and Februacy 1988. All mines had encountered problems with sulphide dust explosions and had introduced defence mechanisms to prevent or limit the extent of these explosions. Techniques included dispersion of limestone by various methods, use of water foggers, placing of water and inert stenuning in blast holes, and removal of settled dust to prevent secondary explosions. The techniques were evaluated and it was concluded that certain practices should be modified or discontinued as they were of little value in preventing dust explosions.

May 11, 1988

By J. P. Sairrlon, K. G. Wheeland, K. J. Mintz, B. J. Kiener

Iwenty-one ore dusts of varying mineralogy, sulphur content and reputed in-mine explosibility were selected from eight mines. After chemical and mineralogical analysis and sieving into two size fractions (+45/ -63 microns and -45 microns), these sarrples and seven pure minerals were tested for explosibility in oxygen in a Harbnann chamber. '!he maximum explosion pressures and maximum rates of pressure do not correlate with sulphur content alone; the types of minerals are also .important. Residues were analyzed to check the combustion reactions. Further comparisons between laboratory and mine results are planned to improve predictability.

May 11, 1988

By Laxman M. Amaratunga

Laboratocy flarmnability tests were perfonned on dust samples from Fraser Mine in a 1.2 L Lindenburg horizontal tube furnace to detennine the percentage of sulphur oxidized during ignition. The explosion pot ential of sulphide dusts depends on their degree of spontaneous oxidat ion, which was measured by the quantitative analysis of so2 gas released during ignition. Test results indicated that dusts with low sulphur content could exhibit a higher degree of oxidation than a sample with high sulphur content. txlst characterisation using image analysis techniques suggests that the structure and the texture of dust are important factors influencing the explosibility of sulphide dust at ignition temperatures. Therefore for safe mine operations% sulphur content cannot be considered as a reliable single index that could be used to predict the potential of a sulphide dust explosion. Physical, chemical and mineralogical characterisation of dust plays an important role in the detennination of suitable indices for the prediction and prevention of sulphide ore dust explosions.

May 11, 1988

By J. D. Botsford

Westmin Resources Limited's H-W orebody on Vancouver Island is a copper/lead/zinc massive sulphide deposit that has been in production since 1985. The sulphur content of the ore averages 28% and in some stopes is as high as 53%. Several minor secondary sulphide dust explosions have occurred in development headings and cut and fill stopes, but as yet no known explosions have occurred in blasthole stopes. This pa:per reviews the history of sulphide dust explosions in the H-W Mine, factors contributing to the explosions and the preventative measures taken to control them.

May 11, 1988

By J. B. McKenzie

Since connnencing underground prcxiuction in 1979, the Ruttan Mine has been prone to secon:iacy sulphide dust explosions. Al though infrequent during the early years of prcxiuction, these occurrences did result in major damage and lost time injuries. 'As the mine exparrled and developed to further depths, sulphide dust explosions occurred on a more frequent basis. Incomplete ventilation se:rvices, required to exhaust the affected areas, caused frequent delays resulting in lower prcxiuction rates due to the time required to ventilate the affected areas of the mine. CUrrent procedures to minimize the effects of secon:iacy dust explosions evolved from practical suggestions arrl examining methcx:ls other mines have developed to cope with the problem. Procedures and regulations to protect errployees from exposure to hazards of sulphur dioxide are time consuming, expensive and counter prcxiuctive in their present fonn. This paper describes the history of secon:iacy sulphide dust explosions at the Ruttan Mine and the methcx:ls and proced.ures currently in place to minimize and contain the effects of dust explosions.

May 11, 1988

By M. Gershon

Four computerized approaches for coal blending are described and their strengths and weaknesses are explored. Spreadsheet analysis Other than word processing, spreadsheet software is the most commonly used software. Its wide availability and ease of use are two desirable features that have led many to try to adapt it to the purpose of coal blending. These are the two main advantages of this approach. To use a spreadsheet for coal blending, each row of data represents one source of coal. Each column represents a quality of the coal, with the last column specifying the tonnage of that coal in the blend. The calculated figures are the blended qualities, which can be programmed to appear in the bottom row. A trial-and-error approach is then used to develop the best blend. Inserting any blend in the last column provides the blend characteristics at the bottom of the page. If any specifications are not met, the blend can be changed, all calculations being immediately updated. This process is repeated until all specifications are satisfied. A third advantage of this method is that it is exactly the procedure that one would go through if doing this by hand. With a spreadsheet, many more blends can be evaluated in much less time, allowing the designer to try to improve over satisfactory blends to find the best of them. The resulting improved blends or cost savings are a fourth advantage of this approach. There is only one major disadvantage of using the spread-sheet. This is that most probably the best blend is not found. As an inexpensive introduction to computer-assisted blending, which can be easily understood and cost just a few hundred dollars, this is a recommended place to begin. Computerized search Computerized search approaches represent crude first attempts at letting the computer design the blend. Unfortunately, the blends found by these programs are very often less satisfactory than those developed manually. The way these techniques work is that they design, according to some predesigned pattern, a number of different blends and then quickly evaluate them. It is then straightforward to select the best from this sample set. The quality of the results depends on the number of blends designed, the way they differ from each other (the search procedure), the number of sources available, the number of quality specifications, and the severity of the specification limits. Good results are usually obtained for small, simple problems. At best, the results obtained are inconsistent from one blend to the next. Programs of this type are advertised in the mining literature. They usually mention how many different blends are evaluated, making it clear that this is more than can be tried manually. Expert systems Expert systems use a source of information known as a knowledge base. The knowledge base, usually a set of if-then rules, attempts to reconstruct the deductive reasoning process of an expert in some job capacity. For coal blending, the usual trial-and-error process has an easily recognized pattern of rules that are followed to try to find a better blend. By following these rules, it is possible to home in on the best blend without the exhaustive search required by the previous method and without the manual manipulation of the spreadsheet approach. The use of an expert system accomplishes this. One of the rules may say something like "if BTU is out of spec, then add more in the blend from a source having a high BTU and subtract some from a source having low BTU." In other words, it goes through the trial-and-error process. The process continues until all specifications are achieved and can continue further to try to optimize the cost. Of course, less simplistic rules are also needed. One set of rules would determine maximum changes in the tonnage added or deleted to the blend or from any one source. Other rules would be needed to determine which quality to improve if two or more are out of spec. At this level, the process of building the knowledge base provides the engineer with a deeper understanding of his or her own thought processes. The advantage of obtaining better blends is clear for this approach. Two more advantages should also be apparent. One is that it follows the same common sense logic that is used when doing the blends by hand or by spreadsheet, making it easy to understand. The other is that, like the search procedure, it designs the blend rather than just analyzing manually designed blends. Thus, it has all the good features of both approaches discussed thus far while providing improved results over each. A final advantage is the ease of use and ease of development. The expert system shells that are on the market are very easy to use and can be run on microcomputers. Linear programming Linear programming is an algebraic tool that solves a set of linear equations to yield an optimal solution to a problem. The optimal solution is based on the linear model so the results are only as good as the ability of the linear model to accurately portray the reality of the problem situation. It is fortunate for the coal blending problem that a linear model can be devel¬oped that reflects exactly the problem that is required. Unfortunately, the mining industry has been slow to use this tool. Coal blending applications available that use linear programming take one of three forms. The first, requiring exper-

Jan 12, 1988

By S. D. Thompson, L. Adler

J.R. Sturgul Thompson and Adler (1988) present a new simulator for use in designing conveyor belt systems for underground mines. Their model uses a large Fortran program to do a Monte Carlo simulation. Their approach is similar to that done by other researchers, such as Stanford (1965) and Newhart (1977). All Fortran based simulation models use an imaginary clock that is advanced a time increment. At each advance of time increment the program has to determine what is to take place in the system. Since many things can take place at different places in the system, this approach requires that a large amount of logic is needed at each step of the program. This, then, necessitates a great amount of jumping between different portions of the program. In fact, the Fortran program, CMBCS, takes up to 15 to 20 minutes to run on the mainframe and several hours for the PC. Such execution times are typical of Fortran based simulation studies. There have been other examples of Fortran based simulation studies in mining. Cross and Williamson (1969) presented a model of a working open-pit copper mine. Fosinato (1982) presented a model of the bucketwheel excavators for the large open-pit mines in Southern Victoria in Australia. These excavators discharge into a complex conveyor system, which can be thought of as being similar, but more complicated than the one modeled by CMBCS. All previous simulation studies using Fortran as the source code have several things in common: they all are very lengthy, often being of the order of 10,000 lines of code and take a great deal of CPU time to execute. The program of Fosinato takes more than 4.5 hours to run on a large mainframe. Also not mentioned by Thompson and Adler is the actual time it took to write the Fortran code. This, too, can be considerable, as Fosinato's program took several man years to write and de-bug. In order to change it, as the parameters change or as the mine increases, further large amount of time is needed. An alternate to such Fortran based programs is provided by using a special computer language known as General Purpose Simulation System (GPSS). This language is designed to solve a wide variety of discrete event simulation problems and appears to be particularly suited for mining situations as shown by Sturgul and Yi (1987). Although GPSS was introduced as far back as 1961, it has not been used in mining to any great extent. However, the original form was quite crude and inefficient compared to modern versions. Perhaps with the increasing popularity and availability of these recent versions, this will change. Two such versions are GPSS/H for the mainframe and GPSS/PC for the personal computer. GPSS/H allows for interactive usage, such as allowing input data to be entered via a menu. GPSS/PC has excellent graphical capabilities and one can actually view the simulation in progress. In addition, one can halt the simulation and change values of various parameters. If a particular portion of the simulation is in progress via a concept known as "microwindows." These new versions have greatly increased the potential for using the language for mining situations. Mining applications of GPSS have been given by Sturgul and Yi (1987), Sturgul and Harrison (1987), Sturgul (1987), and Sturgul and Singhal (1988). Schriber (1987) presents a summary of the recent advances in GPSS.

Jan 12, 1988

By Richard R. Klimpel

Introduction Froth flotation is one of the most common unit engineering operations in use today. Its aim is to upgrade the quality of coal by removing ash/pyrite and separating selected minerals from undesired gangue materials. Flotation has been successfully practiced at the industrial scale for more than 50 years. Its use is increasing due to the ever decreasing feed grades of the coals and minerals being processed. Flotation has also shown itself to be a flexible process that lends itself well to many solid/solid separations. Thus, with surprisingly little equipment modification, separations can involve very different particle sizes and densities, and relative weight ratios of materials to be separated. In addition, the flotation process is economic, especially when compared to size reduction, its associated precursor process. Flotation also lends itself to continuous operations with a given equipment configuration. The operation must exhibit an ability to vary feed rate of solid to the process by as much as 50% without a total collapse of separating efficiency. In addition, the mechanical flotation cell is scaleable from 2.8 to 57 or 85 m3 (100 to 2000 or 3000 cu ft) with surprising ease relative to other unit engineering operations over the same relative size increase. The separating medium used is water. Most of the chemicals required - pH regulators, frothers, collectors, activators, and depressants - are all relatively inexpensive and common. They are not usually used in large quantities. Froth flotation is certainly one of the most versatile and forgiving unit operations in the chemical, mining, and agricultural industries. The problem with the industrial flotation process is that it is fairly easy to get reasonable results and relatively difficult to get excellent results. Why is it so difficult to fine tune the froth flotation process? There are some obvious reasons and a few not-so-obvious factors. The obvious category would certainly contain the tendency of natural feed materials (ores and raw coals) to have variable physical, metallurgical, and surface properties. This occurs even when special care is taken to minimize such differences with blending. There is another obvious reason for difficulty. An inherent broadness of the feed particle size range to flotation results from using any of the common industrial-scalefine grinding devices - ball-, rod-, or autogenous mills. Research has been done into the chemical reagent and equipment aspects of flotation. It has shown strong interactions (limitations) of chemical reagents and equipment on larger and smaller particles. The point is that current flotation practice works best on average- or medium-sized feed particles. The greater the amount of large or small particles, or of both large and small, the more difficult it is to achieve excellent flotation results. Other reasonably obvious factors include: • variations in mineral type and liberation sizes; • water chemistry variations such as those involving Ca", C03, or Fe", or various soluble metal cations; • the presence of stray organic materials such as returning soluble reagents or partially degraded reagents in reuse water; • the lack of reliable on-line instrumentation for monitoring and control; • frequent equipment malfunctions of all types; • excessive feed rates for installed plant capacity; • poor operator interaction with the process; and • poor reagent dosage control. There are less obvious, but real limitations to flotation. These are inherent in the design and operating characteristics of the most common cell design in use today - the mechanical flotation cell. Individual performance attributes of the chemical reagents used, along with their manner of use, also carry limitations. The bulk of this paper will deal with chemical reagents and their use. Chemical reagents In 1979, the author and his co-workers were asked to invent and implement, at the industrial scale, new flotation reagent chemistry. After reviewing some of the literature (Sutherland and Wark, 1955; Klassen and Mokrousov, 1959; M. Fuerstenau, 1962; D. Fuerstenau, 1976; and Leja, 1982), a series of carefully controlled plant scale tests were set up. These tests were designed to understand existing reagent scale-up from the laboratory to the plant. They were to quantify at the plant level the effect of some of the more important, controllable chemical reagent factors (frother and collector type, frother and collector dosage, pH effects). This is a learning experience that all researchers involved in flotation theory should experience. It quickly became apparent that: • because a factor is important

Jan 12, 1988

Power cable costs are only a small part of total mining costs. So many mine operators consider power cable failure and resultant downtime as part of the cost of doing business. But, viewed in terms of lost production, these costs can be quite significant. Now one company, Cablec, seeks to cut cable costs by upgrading the polymer compounding process used to make cable insulating and semiconducting materials. Cablec is the leading manufacturer of electrical power cables in North America. And with about a third of the market, Cablec is the largest supplier of power cable to the mining industry in the United States. To improve its products, Cable has entered the polymer compounding business. In July, it began producing insulator and semiconductor polymer compounds at its plant in Indianapolis, IN. "This new facility provides a quantum leap over conventional compounding methods," said Harry C. Schell, Cablec's president and chief executive officer. "The Cablec polymers plant is producing a dramatically higher standard of polymer compounds that provide significantly higher levels of performance and improved life cycle costs for power cable." Cablec faces tough foreign competition in the wire and cable business. Competing on price alone is difficult, particularly when foreign producers are state subsidized. So Cablec feels the best way to compete is to establish new quality production standards. The company's new polymers plant is one way to do this. By increasing purity control and uniformity in polymer compounding, Cablec says its power cables will last longer and fail less often. A typical medium voltage cable consists of a conductor, conductor shield, insulation, insulation shield, metal shield, and jacket. The conductor shield and the insulation shield are conducting polymers. Contaminants and imperfections can occur within the insulation, at the conductor shield/insulation interface, or at the insulation shield/ insulation interface. Over time, these contaminants and imperfections can decrease the electrical strength of the cable or cause premature cable failure. The effort to minimize the number and size of any possible contaminants begins with pure polymer compounds mixed in a clean facility. However, most power cable manufacturers manually handle raw materials, use ethylene/propylene (EP) in bulk bales, and mix polymercompounds in open Banbury mixers. The quality and uniformity of polymer compounds is also impacted by temperature variations in the mixing process. This results in wide gradations of product consistency from batch to batch and ultimately contributes to power cable failure. Cablec says the improved polymer compounds from its state-of-the-art plant will be the purest and most consistent insulating and semiconducting materials available. The plant itself RCA spent $18 million to build Cablec's Indianapolis plant. RCA used the facility to mix specialty polymer compounds used to make video disks. RCA had two considerations in mind for the plant, cleanliness and uniformity of the compounds. However, when the video disk market failed to materialize, RCA sold the 46.5 dam 2 (50,000 sq ft) plant to Cablec for $3.1 million. Cablec invested an additional $3 million for modifications and increased production capabilities. Today's replacement cost for such a facility is estimated at $30 million. Cablec says the plant will set a new standard for performance and be economically difficult to duplicate anywhere. One of the essential elements of the plant's clean process environment is the air intake system. It filters contaminants greater than 2 um, less than one-fiftieth the current industry standard. All material handling and conveying areas in the facility are air-locked. This keeps out contaminants such as smoke, dust, and pollen. Banks of pneumatic pumps move polymer components through the system and continually filter the air. The plant also has a backup air intake system. No process downtime due to pump failure here. From the time raw material enters the plant, it is stored, transported, and processed in filtered air by an airtight stainless steel system. The stainless steel resists rust and corrosion. This further eliminates the danger of contamination from paint or rust particles in the conveyance network. A computer system allows a single operator in a central control room to monitor every aspect of the compounding process from air quality to line speed. The computer

Jan 12, 1988

By S. Hodgson

Introduction The Red Dog zinc/lead deposit in northwest Alaska is the second largest zinc deposit ever discovered. Being developed by Cominco Alaska, with startup scheduled for early 1990, it will ultimately be the largest zinc mine in the Western World. Table 1 shows the reserves of the deposit. Bob Baker, a local bush pilot, discovered Red Dog in 1968. The US Geological Survey assessed the find. But the Red Dog site had been withdrawn from staking by provisions in the Alaska Native Claims Settlement Act of 1971 (ANCSA). Under the terms of ANCSA, the deposit was selected by the NANA Regional Corporation, one of 13 native corporations established by the Act. Then, in 1980, the US Congress confirmed this selection, in the Alaska National Interest Lands Conservation Act. In 1982, Cominco and NANA reached an agreement that has led to the development of the mine. Under the agreement, Cominco leases the property from NANA, develops and operates the mine, and markets the concentrate produced. In return, NANA receives royalties and, after capital recovery, participates in the net proceeds from the operation. The Alaska State Government is also playing a role in the development of the mine. Under the auspices of the Alaska Industrial Development Authority (AIDA), the state will construct, own, and operate the DeLong Mountains Regional Transportation System. This consists of a 84-km (52¬mile) road from the minesite to the Chukchi Sea. Located here are the port facilities necessary to store and ship the annual concentrate production and receive the mine fuel and supplies. For using this transportation system, Cominco will pay an annual toll fee. Over the anticipated 50 year life of the operation, the fee will provide a 6.5% return to the State of Alaska. In late 1986, Cominco's Board of Directors approved the Red Dog project. And, during the summer of 1987, road construction began. In February 1988, initial stripping began at the ore body and construction started at the millsite. Project setting The project site is approximately 1000 km (600 miles) northwest of Anchorage, 84 km (52 miles) east of tidewater on the Chukchi Sea, and 192 km (120 miles) north of the Arctic Circle (Fig. 1). The countryside consists of low, rolling hills that lead into mountains north and east of the deposit. The vegetation is typical tundra mat. And the site is in an area of permafrost. However, the permafrost is relatively warm and some areas have been found to be thawed. Examples would be creek bottoms and the site of the main waste dump.

Jan 12, 1988

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. B. Jarboe, H. Wen

Introduction The 150 MW circulating fluidized bed (CFB) power plant described in this article is the product of recent Bechtel experience in the industrial power and steam market, where the economic feasibility of fluidized bed projects has been proven. The ability of CFB boilers to burn low grade fuels makes it possible to burn low ranking coals or coal wastes. These include anthracite culm, coal mine rejects, and coal wash-plant-waste slurries. This article also addresses the impact of fuel characteristics on plant design. This includes solids handling, steam generation, sulfur capture, and auxiliary load requirements. The uncertain economic environment in which utilities must now operate has shifted the industry's interest away from large, base-loaded generating units to small, more versatile installations. Utilities now consider purchasing new generating capacity in smaller increments and staging its deployment to correspond with demand increases. In this way, they lower their risk in the increasingly hazardous art of long-range load forecasting. This strategy has many advantages. Multiple, smaller units have inherently greater turndown capability and better availability than a single larger unit of the same size. In addition, the smaller units will each have better cycling capability. And, with CFB combustion, the smaller units will have greater fuel flexibility. General plant description The 150 MW powerplant described here is a grassroots CFB unit. Fluidized bed technology differs from conventional combustion techniques. With CFB, solid fuel, together with inert material such as sand, gypsum, limestone, or ash from the fuel, is kept suspended in the combustion chamber. This is done by using the action of the fluidizing air distributed below the bed. Fluidization promotes the turbulent mixing conditions required for good combustion. The turbulence causes the entire mass of solids to behave like a liquid. The result is improved mixing. This allows heat to be generated at a lower and more uniformly distributed temperature that is below the ash fusion temperature of most coals. The fluidizing mechanism offers several other advantages over conventional combustion. There is less volatilization of alkali compounds, which cause deposits on boiler tube surfaces. And there is less sensitivity to the quantity and nature of the ash in the fuel. The low combustion temperatures prevent vitrification of the ash particles and minimize problems associated with furnace slagging. The ability to select from many fuels is one of the major advantages of CFB combustor technology. Fuels that can be burned in CFB plants to produce steam and electricity include: low grade coals with high ash or sulfur, lignite, peat, wood and pulp products, biomass, refuse, anthracite culm, petroleum coke, and oil shales. The plant's turbine cycle uses superheated steam at 12.4 MPa (1800 psi) and 538°C (1000°F), with single reheat to 538°C (1000°F). Two high pressure and three low pressure, closed, feedwater heaters are furnished. So is a deaerator, which is an open, direct-contact heater. The CFB plant exceeds EPA's New Source Performance Standards (NSPS) for air emissions. The design limits sulfur dioxide (SO,) emissions to 0.25 kg/GJ (0.6 lb per million Btu). SO2 reduction occurs in the combustor by reac-

Jan 11, 1988

By Kelly F. Bailey

Introduction On June 20, 1986, the Occupational Safety and Health Administration (OSHA) published revised asbestos exposure standards for general industry and construction. The standards reflect OSHA's attempt to adequately control workplace exposures to minerals it considers carcinogenic - minerals capable of causing or contributing to cancer. These standards specifically identify asbestos as: chrysotile, an asbestiform serpentine mineral; and the amphibole minerals amosite, crocidolite, tremolite asbestos, actinolite asbestos, and anthophyllite asbestos. Each of these has a more common nonasbestos mineral analog that exists in nature in a crystalline, blocky shape rather than the hair-like or fibrous shape of asbestos. The mineralogical names for three of these nonasbestos minerals are unique: antigorite for chrysotile, cummingtonite-grunerite for amosite, and riebeckite for crocidolite. The other three nonasbestos analogs do not have unique mineralogical names. They are simply designated as actinolite, tremolite, and anthophyllite without the word asbestos following their names. The 1986 OSHA standards not only cover exposure to the six asbestos minerals, they also cover specifically the nonasbestos forms of actinolite, tremolite, and anthophyllite (AT&A). The new standards regulate these minerals exactly like asbestos (OSHA, 1986). The construction aggregate industry views this as a major problem because these nonasbestos minerals are common amphibole rock-forming minerals in the earth's crust. They exist in small quantities over large areas of the United States (Kuryvial et al., 1974). These minerals, unlike asbestos, are not mined for a specific commercial purpose. They are unavoidable components in much of the aggregate used for construction throughout the US. They are also common in the gangue material of metallic ores. There are areas of the US where amphibole-bearing bed¬rock is common. Not every rock mass in these areas contain amphiboles, however. It does mean, though, that amphiboles are physically compatible with many of the rocks in those areas. And given the correct geochemical conditions, they will be present primarily in the nonasbestiform variety. In addition, these amphiboles will probably exist in the natural drainage system, sand and gravel deposits, stream sediments, lake shores, valley basins, or ordinary beach sand within these areas. There has been little quantification of nonasbestiform AT&A in dusts and soils in the US. This is not surprising since these nonasbestiform minerals are not commercially valuable. However, an example of the pervasive nature of these minerals can be found in a 1981 Geological Society of America publication where about 0.7% tremolite-actinolite was found in the desert dust in and around Tempe, AZ (Pewe, 1981). Since OSHA standards treat these common nonasbestos minerals as carcinogens in the same way as asbestos, large natural areas in the US are implicitly being labeled as hazardous by OSHA. When a substance is identified as a carcinogen, another OSHA standard comes into play, the Hazard Communication standard. There are also right-to-know laws in 9 states that essentially duplicate this federal standard. These standards require that a product containing 0.1% or more of an OSHA-designated carcinogen be labeled as such (OSHA, 1983). This means that much of the stone and sand gravel products occurring naturally and mined in the US could be labeled a carcinogen when, in fact, they are not. The National Stone Association (NSA) and the domestic construction and mining industries believe that OSHA has seriously erred. The NSA has studied the health, mineralogical, technical, economic, and legal basis for OSHA's action. These studies concluded that there is no justification for the agency regulating nonasbestos minerals as if they were asbestos. Health issues The preamble to OSHA's 1986 asbestos standard states that evidence for asbestos-like health effects from exposure to nonasbestiform varieties of AT&A is inconclusive (OSHA, 1986). The fact is, not only are the data inconclusive, they are nonexistent. During 1986-1987, NSA's occupational health and epidemiology consultant, Environmental Health Associates (EHA), reviewed all available health studies related to AT&A. EHA found evidence that malignancies in both experimental animals and humans are associated with the asbestos forms of these minerals. No experimental or epidemiological evidence was found that indicated such pathogenic effects occur from exposure to nonasbestiform varieties of these minerals. There are relatively few scientific studies of the health effects of exposure to nonasbestiform varieties of AT&A. In three different animal studies, exposure to either nonasbestiform tremolite or actinolite did not result in pulmonary fibrosis on in excess tu-

Jan 11, 1988

By Tim Neil, O&apos

Finnish mining traditions date back 400 years. The country was then one of Europe's major iron producers. Most of the mines in current operation, however, were started during the last 20 to 30 years. The Finnish mining industry must make do with limited resources. Deposits are usually low grade and small in size. Open-pit methods are used where possible. But thin ore bodies and harsh winters can impose severe restrictions. Underground mining is often the only option. Despite these handicaps, Finnish metal miners produce copper, nickel, zinc, vanadium, iron ore, cobalt, and ferrochrome. Ironically, the success of Finnish mining is rooted in a crisis the industry faced several decades ago. Mining costs were rising. High grade mineral deposits were nearly exhausted. It was vital that Finland develop new mining machinery and methods that afforded higher productivity. Cooperation between the Finnish mining industry and its equipment manufacturers was good. Partly as a result, the Finnish mining industry has turned itself into one of the most modern and efficient in the world. The Finns did this while continuing to meet the stringent safety and environmental requirements in force in the Nordic countries. Today, the share of domestic raw materials used by the Finnish metallurgical industry is decreasing. So the country increasingly depends on imports of ores and concentrates. To better provide these imports, Finnish mining companies have expanded their foreign operations. And, while importing raw materials, the mining industry has been exporting its technological and equipment know-how and expertise. Mining methods Conventional mining methods and machinery are used in surface mines. Drilling is the exception. It is done with Tamrock hydraulic track drills and 100 to 150 mm-diam (4 to 6-in.-diam) drill holes. Pumpable slurry explosives are used in larger open pits. Underground, sublevel stoping with open stopes is the most common mining method. It accounts for about 80% of all underground stoping. The main benefits of this method are low production costs, high mechanization, and easy control. The best productivity figures exceed 10 t (11 st) per man-hour, including all mining personnel. The development of cable bolting has made it possible to use sublevel open stoping even in relatively weak rock. Cut-and-fill methods, such as concrete-pillar stoping, inclined-wall stoping, and traditional cut-and-fill have also been used, combined with sublevel caving and room-and-pillar methods, according to local demands. Various fill materials and stabilization components have been investigated and tested for cut-and-fill methods. Sublevel open stopes are often later filled, to stabilize the surface or to allow pillar recovery.

Jan 10, 1988