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By G. H. Matheson, J. M. W. Mackenzie

The coal concentrates studied contained clay slimes which would not flocculate on the addition of a polymer flocculent alone. The mechanism of the flocculation process is interpreted on the basis of the flocculation of lyophobic colloids. The filtration rate characteristics are shown to deviate slightly from certain basic filtration equations. Cake moisture was shown to decrease with cake drying time and vacuum, and increase with cake thickness. Filtration is a commonly accepted method of dewa-tering the fine coal concentrates produced in coal cleaning operations. Dewatering of these concentrates, using drum or disc vacuum filters, usually presents few technical difficulties and will result in the production of a filter cake containing some 20 to 25 pct moisture. In many coal cleaning operations it is common practice to add flocculants to the coal concentrates prior to filtration to increase the filtration rate. The mechanism whereby flocculation increases filtration rate has been summarized by Gaudin,4 while Geer, et a15 have compared the effectiveness of starch and synthetic flocculants as filter aids for a variety of coal concentrates. Coal concentrates, containing appreciable quantities of expanding lattice clays of the montmorillonite-illite group, sometimes prove impossible to filter without flocculation of the pulp prior to filtration. The fine coal concentrates produced at the Newcastle washery of The Broken Hill Proprietary Co. Ltd. fall into this group. Coals mined in this district contain appreciable quantities of expanding lattice clays which break down in the washery circuit waters and, by processes of mechanical inclusion, contaminate the coal concentrate thickener feed. Despite the classifying action of this thickener, the underflow which is pumped to the filters contains sufficient clay slimes to render filtration of the unflocculated pulp impossible. Filtration leaf and plant tests were carried out to develop effective and economic flocculant combinations and to determine the filtration characteristics of the flocculated coal concentrates. The determination, by leaf testing, of the filtration characteristics, and the analysis of these characteristics in the light of the equations of Brown2 resulted in the successful trial operation of the filters in the Newcastle washq. Although the filtration characteristics are not entirely in agreement with the theoretical treatment of Brown, the equations developed made it possible to predict, from leaf tests, the effect in the plant filters of pulp density, form vacuum, and form time on filtration rate and filtrate volume. EXPERIMENTAL PROCEDURES The sample used for leaf testing was the coal concentrate thickener underflow, which in plant operations would form the feed to the two 12-disk Dorr-Oliver filters. The feed to the coal concentrate thickener comprises the undersize produced by dewatering the Diester table concentrate on 5-mm dewatering screens and the overflow produced by cycloning of the froth flotation concentrate. At the time of testing these filters were inoperative because of the difficulties in filtration, arising from contamination of the

Jan 1, 1962

By G. A. Vissac

The drying of fine coal involves special techniques, which are discussed and analyzed. Types of dryers employing these techniques are described. Calculations are presented for new methods of dealing with the entrained dust that is always present in fine coal drying operations. NEW conditions, new requirements, and new methods have increased the demand for more efficient and more economical methods of drying fine coal. Dewatering of larger sizes may reduce the surface moisture to 8 or 9 pct. It is more difficult, however, to dewater sizes below 1/4 in., and some filter cakes still contain as much as 20 or 25 pct moisture. Increased freight rates and stricter consumer specifications have resulted in a demand for further reductions in moisture content. This can be obtained only by heat drying. Most modern methods of heat drying disperse or spread the mass of coal to be dried, in an atmosphere of dry hot gases. The more intimate the contact between coal particles and hot gases, the quicker and more efficient the drying operation will be. Two different techniques are generally employed, using either a fluidized condition or an entrained condition of the coal to be dried. Fluidized Condition Fluidization of a body of sand was defined and explained by Fraser and Yancey in a paper published in 1926.' This condition was artificially obtained and maintained by proper regulation of the rate of air flowing through the sand body. "The sand bath 'boils' uniformly on the surface," they write, "and feels like a fluid." The fluidization technique was also described and analyzed by Steinmetzer2 in connection with the operation of an air cleaning table. His main conclusions are as follows: "Fluidity is, for the particles involved, the possibility of motion with minimum friction. . . . Then fluidity requires the introduction of various forms of energy capable of neutralising frictions. Two solutions can be used— air and/or mechanical motions (such as the shaking motion of the carrying deck of the air table). The combination of mechanical and air energy will give the widest margins of size ratios and of bed thickness, translated in capacity per unit area of the carrying table." Richardson and Langston3 have indicated results obtained with a dryer working with a fluidized bed. They used a vertical tube type of dryer, however, without the assistance of any mechanical energy, and without any lateral motion of the fluidized bed. The capacity of such a dryer is too limited for practical applications, since the speed of the acceptable air currents is held to the speed of fall of the particles involved. Capacities as low as 182 Ib of coal per hr per sq ft of dryer area are indicated. As stated by Richardson: "A basic limitation to a fluidised bed dryer is that the velocities of the gas must be held within a definite range; with velocities of 10 ft per second, all coal minus 6 mesh in size will be entrained, and the operation is then similar to that of a Flash dryer." A fluidized bed must be virtually static. The coal particles simply kept in suspension offer a minimum resistance to the flow of gases, insuring the most favorable conditions for rapid evaporation of surface moisture. However, very wet fine coal, i.e., over 12 pct of surface moisture, will be delivered in the forms of mud balls, or as a soggy, sticky mass, almost impossible to disperse, sticking and acting as a wet blanket on the deck. Strong currents of gases and wide deck perforations will be required to punch holes in the wet mass and gradually loosen and fluidize it. The mechanics of fluidizing a bed of coal in a gas medium for the purpose of obtaining the most efficient drying condition are entirely similar when the fluid used is water and the purpose is to break up and distend a bed of coal to be cleaned so that perfect stratification according to densities will be insured. Purely mechanical energy is used in the basket-type jig, water pulsations in the piston and in the Baum-type jigs. A combination of mechanical motion and of air pulsation offers the most efficient and favorable conditions. Entrained Condition The most critical factor to be considered in the design of a dryer employing the entrained condition technique is the speed of the hot gases to be circulated in the drying column. With insufficient gas velocity, excessive amounts of the largest sizes will drop to the bottom of the dryer column without being thoroughly dried. On the other hand, high gas velocity will cause degradation, dust losses, and high power consumption. Figs. 1 and 2, reproduced from Hanot,4 show the relative importance of speed and temperature for various sizes of particles. It can be seen, for instance, that to maintain in unstable equilibrium particles of 1/4-in. size in a gas current at 500°C, a speed of 30 meters per sec, or 6000 fpm, will be required. For % -in. particles an almost prohibitive speed of 45 meters per sec, or 9000 fpm, will be necessary. In practice, maximum gas velocities of 3000 fpm are recommended; since power increases as the cube of the velocity, it can be seen that beyond certain limits such dryers would not be economical. If the particles were moving at the same speed as the hot gases they would remain in the same

Jan 1, 1954

By E. W. Gieseke

Coal preparation men today are confronted with problems brought about by changing feed characteristics. The size consist of the coal going to preparation plants has been getting finer and finer. In order to handle these sizes economically, operators are turning to Notation. American Cyanamid Co., in an effort to enable coal flotation plants to produce efficiently at maximum quality and capacity, wnducted experiments which demonstrated that synthetic flocculants are useful for increasing the capacity of coal and refuse filters and for clarifying coal and refuse slurries, and a wetting agent is beneficial in reducing the moisture content of filter cake. Coal preparation men are today confronted with problems brought about by the changing coal markets, changing feed characteristics, and ever increasing labor and supply costs. Small mines are finding it difficult to exist, and the larger mines are rapidly converting to fully mechanized and automated operation to survive. The size consist of the coal going to preparation plants has been getting finer and finer, and the problems facing the preparation plant personnel are growing larger and larger. Facilities can be provided to handle this plant feed in new plants, but the problems are generally not so simple in existing plants where equipment must often be used to perform tasks far beyond its rated capacity. In addition, various state conservation departments, water commission boards, and other Federal and state agencies are endeavoring to eliminate stream pollution. These problems, and many more not mentioned, have been forced on the preparation engineer for reasons of economy and the public interest. In the matter of stream pollution, a determined effort is being made by the coal industry to cooperate with the Federal, state, and local agencies. Economists predict it will be necessary to produce 600 million to 700 million tons of coal by 1970. Bear in mind that the greater proportion of this increased tonnage will have to be processed in existing plants, and you get an idea as to the problems facing the preparation engineer. Methods presently in use for preparing fine coal do an acceptable job down to 48 to 100 mesh. A very large percentage of the - 48 or -100 mesh fraction of the tonnage of coal produced now is being sent to settling ponds or sold at a low figure for reason of quality. This has caused operators to turn to flotation. Building settling ponds and impounding fines is a costly and, at times, impossible method of handling these sizes. Recently a number of flotation plants have been installed to beneficiate material finer than 48 mesh, and these plants have demonstrated that flotation can be successful and increase the plant realization. Coal is a low cost commodity and processes for its beneficiation must be low in cost. Flotation is such a process. It is easily adapted to handling fine coal, requires very little supervision when properly installed, and is economical to operate. Flotation separates coal from refuse and fine clay. For reagents, several combinations are used. Some

Jan 1, 1962

By J. P. Morgan, J. R. Dawson

A review of Jones & Laughlin's Vesta coal preparation plant flowsheet in 1959 indicated that, of the three sizes making up the company's metallurgical product, the -48-mesh fraction offered the greatest potential for improvement. Efforts to improve quality and yield were directed toward the use of froth flotation. After laboratory and plant tests were completed, it was decided that treatment of the entire -@-mesh fraction of the coal mix by flotation was a significant step forward in providing better coal quality and recovery. Reduction of ash in the metallurgical coal from 6.75 to 6.25% provided material savings at the coal washer, coke plant and blast furnace. The Vesta preparation plant of the Jones & Laughlin Steel Corp. is located on the Monongahela River, eight miles south of Brownsville, Pa. This plant is capable of cleaning 2400 tph of a mixture of Pittsburgh seam high volatile coals from the Vesta and Shannopin mines to produce a metallurgical coal of 6.75% ash and a steam coal containing 14% ash. Heavy media vessels are utilized to upgrade the coarse coal while launders and tables handle the -1/4-in. fraction of the raw coal feed. Cleaned -1/4-in. coal goes to boot classifiers where the 48 mesh material settles and is removed by bucket elevators. This 1/4x48 mesh material is then sent to centrifugal dryers for dewatering. As designed, the -48 mesh fraction overflows the boot as a 7 to 8% solids slurry and is recovered by cyclones and a thickener and then dewatered by filtration. The products of the heavy media circuit, the launder-table-boot-dryer circuit and the cyclone-thickener-filter circuit, are combined to constitute the shipped metallurgical coal. The Vesta-Shannopin mix is combined at Jones & Laughlin's Pittsburgh and Aliquippa Works with purchased low and medium volatile coals for the production of metallurgical coke. A review of the preparation plant flowsheet in 1959 indicated that of the three sizes making up the metallurgical product, the -48-mesh fraction, containing 16.7% ash, offered the greatest potential for improvement of coal quality. This fraction was considered low, averaging 55%, while the rejects averaged only 33% ash. It was evident that while the boot overflow fraction represented only 6% of the product mix, a reduction in ash to 6 or 7% would reduce the ash of the metallurgical coal by 0.5%. The original plant flowsheet handling the boot overflow (Fig. 1) consisted of ten 14-in. cyclones treating 70% of the overflow and producing a 48x150-mesh underflow and a -150-mesh overflow. The cyclone overflow was combined with the remaining 30% of the boot overflow to constitute the feed to a 150-ft thickener. The cyclone underflow and the 48x325-mesh thickener underflow were fed to vacuum disk filters for dewatering prior to combination with the coarse metallurgical coal. The thickener overflow stream containing the -325-mesh material was split, 40% being recycled directly to the plant water system and 60% being flocculated and clarified in an 85-ft thickener. The underflow of this thickener was pumped to waste. Efforts to improve quality and yield on the -48-mesh coal were directed toward the use of froth flotation early in the investigation. A review of the fine coal flowsheet indicated that three streams should be considered for treatment by froth flotation. These were the -325-mesh thickener overflow, the -150-mesh cyclone overflow and the -48-mesh boot overflow. Flotation of the thickener overflow was considered basically to improve recovery by upgrading the reject stream so that it could be fed to the filters along with the thickener and cyclone underflows. The overflow from the thickener contained 3.7% solids and had an ash analysis of 33%. Flotation of the cyclone overflow and boot overflow were considered to upgrade coal quality as well as yield. Treatment of the cyclone overflow, a 5.4% solids slurry averaging 27% ash, would eliminate the coal thickener and provide a feed of flotation concentrate and cyclone underflow to the filters. Treatment of the boot overflow would eliminate both the cyclones and thickener and feed only flotation concentrates to the filters. Flotation feed in this

Jan 1, 1962

By K. I. Savage, S. C. Sun

This paper describes a process developed to recover coal, clays and pyrite from coal wastes. The process consists of fine grinding followed by coal and pyrite flotation which leaves the clays in the flotation pulp. A bituminous coal refuse containing 10% sulfur and 30% carbonaceous material was treated by this method to yield a coal product containing 4% sulfur, 10% ash; a pyrite product containing 45% sulfur (84% FeS2), 1% carbonaceous material; and a clay product containing 2% sulfur (3.5% FeS2). The coal yield was about 89%. The pyrite yield was about 77%. The process steps may be entirely flotation, or gravity separation (hydrocycloning) may be used to increase the pyrite : coal ratio in the flotation feed. Cost estimates for the process show a profit of $2.28 per ton of low pyrite grade refuse, but these do not include labor, maintenance, overhead and plant depreciation. The development of this process consisted of three parts: (1) exploratory tests, (2) op-timazation tests and (3) confirmatory tests. The objectionable qualities that sulfur imparts to coal have been commented upon from early times, and they have become more objectionable as the uses of coal have grown. In whatever form coal takes — raw, carbonized or gasified —the sulfur content remains objectionable, and therefore its compounds are removed as completely as possible. It has long been known that sulfur occurs in coal in different forms. In 1861, sulfur was said to exist in the state of sulfuric acid in combination with a base; in combination with iron as iron pyrites; as bisulfides of iron; and in combination with the organic elements of coal.' Pyritic sulfur is a term loosely used to cover the sulfur associated with iron in its various forms. The mineralogically recognized forms are pyrite (FeS2), pyrrhotite (Fe 1-x S) and marcasite (FeS2). The particle sizes of pyrites vary widely. Isolated grains of marcasite smaller than 15 microns have been found disseminated through coal.2 Hair-like "veins" of pyrite filling cracks in vitrain have been found. At the other extreme, lumps or nodules of pyrites large enough for removal by hand picking have been encountered. Organic sulfur, unlike pyritic sulfur, does not exist as discrete particles, but is instead intimately associated with the coal structure and thus it is impossible to remove it or reduce its concentration by physical or mechanical means.2 In the preparation of coal for its various markets, the pyrite minerals (pyrites) are separated from the raw coal feed. This separation process concentrates the pyrites in the tailings or other waste products. It would be desirable to recover these pyrites for three reasons: (1) Pyrites are potential sources of sulfur and iron. In 1967, for the fifth consecutive year, Free World consumption of sulfur exceeded production.3 Propelled by the shortage, the domestic price of sulfur has risen from $24 to $38 per long ton (bright). (2) The refuse, when placed in piles, becomes ignited. The pyrites (FeS2) bum, giving off SO,. Thus, the pyrites are a cause of air pollution. (3) Also, the pyrites undergo chemical reaction when exposed to air. The refuse is leached by waters which result in stream pollution due to the water-soluble iron and acidic reaction products. Coal refuse also contains coal minerals and clay minerals. Therefore, any process for recovering the pyrites must successfully separate them from the coal and clay minerals. In the study discussed here, ten different bituminous coal refuse samples were successfully upgraded in pyrite content. These samples represented a wide variety of coal waste materials from Pennsylvania and other states. The variabilities of the sulfur and coal contents are shown in Table I. The extremes are Sample J, a high sulfur-low coal, and Sample B, a low sulfur-high coal. Definitions of some of the symbols or terms used in this report are given below: Mesh size—Tyler standard mesh screen sieves with an opening based on the square root of two. Fe —Indicates iron, as determined by a stannous reduction-dichromatic oxidation method. S—Indicates sulfur, as determined by the ASTM "Eschka" method for sulfur in coal.

Jan 1, 1969

By J. D. Allen

In the coal preparation plant at U.S. Steel Corp.'s Concord mine in Alabama, the product moisture was reduced from about 20% to 9% through the development of a workable flowsheet for dying the clean coal. Some of the high points of the conversion of the Concord washer from vibrating screen dewatering to centrifuge drying are presented in this article. The coal preparation plant at the United States Steel Corp.'s Concord mine near Bessemer, Ala., began operations in June, 1950. The basic cleaning circuit consisted of a Chance cone for cleaning the 2-in. x 3/8-in. fraction and concentrating tables for cleaning the -3/8-in. coal fraction. Dewatering of the final product was accomplished with 35-mesh vibrating screens. However, because Concord coal is very friable with 85% passing 3 mesh and 25% passing 28 mesh (Fig. I), a considerable quantity of fines passed through the dewatering screens. To recover these fines, the dewatering screen effluent was sent to two 70-ft-diam hyd rose para tors from which the underflow returned to the dewatering screens and the overflow went to a 175-ft-diam thickener for water clarification, as illustrated in Fig. 2. As might be expected, there was considerable recirculation of the -35-mesh coal in this system. This degraded the coal and led to higher moistures. It also resulted in a fines bleed-off of plant water and thickener underflow to a fine-coal settling pond. In 1961, vacuum filters were installed to supplement the screens and to recover the fine coal lost in the bleed-off. The above system resulted in a product moisture of about 20%. Further reduction to approximately 11% was accomplished by drainage from railroad cars during transit to the coke plant. Control of the moisture at unloading was poor because of variations in operating levels and railroad car availability, but since the Concord coal constituted a relatively small percentage of the over-all coal mix at the coke plant, its high moisture and fluctuations did not seriously hamper coking operations. Early in 1964, however, the proportion of Concord coal was increased with a resulting increase in the moisture of the total coal mix, and this caused difficulties in the coke plant. The increased participation of Concord coal also brought about a shortage of railroad cars which raised the moisture content to a point where the coal became almost impossible to handle. In an effort to alleviate this problem, U.S. Steel initiated a research program to develop a flowsheet for reducing the moisture to approximately 10% and to determine whether the resulting product would be satisfactory for the coke plant. Under this program, an experimental flowsheet for evaluating both a vibratory centrifuge system and a solid bowl centrifuge system was drawn up as shown in Fig. 3. A horizontal vibratory centrifuge was installed to reduce the moisture in the dewatered coarse coal, and a four-cell, flotation machine was installed to clean the -65-mesh portion of the 3/8-in. x 0 table raw coal feed. Several shortcomings in the above circuits were apparent. For one thing, large quantities of repulping water were required. Furthermore, since about half of the flotation feed was of a size consist that could have been cleaned on the tables, these circuits would result in an unnecessary increase in reagent costs and facilities. Maintenance of the hydro separators and pumps for pumping the underflow to the filters,

Jan 1, 1969

By Harold R. Bradford

EXPANSION initiated during and after the war has placed industrial plants in new areas and increased reduction and manufacturing facilities in communities already established. With added expansion interest has grown in possible area contamination from waste products. Evolved flue waste contains many chemical contaminants that may or may not cause concern; however, atmospheric pollution has always been a problem to industrial communities. For some time effects attributed to small quantities of compounds of fluorine in various waste gases have been studied by industrial, medical, agricultural, and associated interests.1'2 Aside from an article by Churchill, Rowley, and Martin in 1948," little has been published on the subject of fluorine in coal except by foreign investigators. The first indication of fluorine compounds in coal was noted in England in 1934 by R. Lessing,4,5 who found that the presence of fluorine in Midland and West County coals was definitely established by etching of glass and by severe corrosion of porcelain fillings in a gas works scrubber. Etching of glass in a direct contact annealing kiln initiated research, and in 1944 four papers were published by H. E. Crossley,6-9 giving results of investigations of fluorine in coal of England by the Fuel Research Board, Department of Scientific and Industrial Research. These investigations studied methods of determining fluorine, the quantities present in various coal deposits, the geographical distribution of fluorine in British coals, occurrence of fluorine in minerals associated with coal, amounts of fluorine compared to other constituents of coal, the nature of fluorine compounds, and industrial significance of fluorine in coal. It was shown that British coals contain from 5 to 200 ppm of fluorine, which is probably present as fluorapatite [Ca10F2(PO4)6], often written as [3Ca3P2O8.CaF2]. The fluorapatite is intimately associated with shale in some coals and can often be separated from these coals along with the shales. H. E. Crossley8 also reported no correlation between the amounts of chlorine, sulfur, or ash and the amount of fluorine in different coals. He found, however, a relationship between the content of fluorine and phosporus in many British coals. Thus it should be possible to anticipate likely fluorine-rich coals from existing data giving phosphorus content. Behavior of the fluorine compounds of coal was compared with that of 12 fluoride-phosphate minerals in sink-float separations and solvent ex-

Jan 1, 1958

By C. F. Cockrell, K. K. Humphreys, H. E. Shajer

It is quite conceivable that the question of using coal or an alternate fuel in the future by large industrial consumers may be decided on the basis of possible profitable utilization of flyash. Accordingly, the Coal Research Bureau at West Virginia University has been attempting to develop new uses for this waste material. This effort has resulted in the development of a process for producing flyash-based face brick and other structural materials which generally exceed published specifications for competitive clay-based materials. These structural materials, containing up to 97% waste coal ash, can be produced from flyash derived from most coals without regard to the physical or chemical properties of the coal ash. Preliminary production cost estimates indicate that face brick can be manufactured at a rate of 9000 brick per hr or more with a resultant gross profit of over $24.00 per thousand brick. Plant payout time for a plant of this size would be only about 1.2 years assuming a market price of approximately $55.00 per thousand, a well established figure for comparable clay brick. With flyash production expected to more than double in the next 15 years, there is an increasing need for a practical and Profitable method for the utilization of large tonnages of this material (see Table I listing total flyash production of utilities in each state of the U.S.). Present industrial and governmental physical and chemical restrictions, particularly for use as a lightweight filler or pozzolanic material, have limited the use of flyash considerably. For those ashes which have met the required standards, utilization is not always assured because of lack of promotion and poor public acceptance. Less than 6% of the total flyash production in the U.S. was sold in 1962. The remainder was dumped in disposal areas at costs as high as $2.42 per ton.' It is quite conceivable that in the near future the question of using coal or an alternate fuel by large industrial consumers may be decided on the basis of the possible utilization of flyash. To help insure future coal demands, the Coal Research Bureau at

Jan 1, 1968

By Elmer R. Kaiser

COMBUSTION of coal and transfer of heat from flames and gases to boiler surfaces continue to be of great interest to engineers here and abroad. Numerous investigations have been in progress to improve furnace and boiler performance and economy. The importance of better understanding of the processes and opportunities for improvement is apparent when it is remembered that heat from at least 500 million tons of coal a year the world over is being transferred to boiler water at efficiencies ranging mostly between 50 and 90 pct. Even slight gains in efficiency, economy, and labor saving become very significant when multiplied by the enormous quantity of fuel consumed. Also the competitive position of the large coal, oil, and gas industries in satisfying the fuel consumers is greatly affected by the achievements made through technical progress with each fuel. This paper is part of a continuing activity of Bituminous Coal Research, Inc., to extend the knowledge of coal utilization for steam generation and to seek promising directions for future research and development in cooperation with others. Particularly in the latter regard, numerous interviews were held during the last three years to seek the experience and advice of boiler and combustion-equipment manufacturers, electric-utility executives, and fuel engineers. A wealth of published information was also reviewed, which together with the interviews pointed to the advisability of further work on ash and sulphur control. For the present purpose a number of factors important to efficient heat liberation and recovery have been grouped as follows: 1—combustion, temperatures, and rates of heat liberation; 2—radiation, convection, and furnace and boiler configuration; 3—sponge ash, slag, and hard-bonded deposits; 4— low-temperature deposits and corrosion (cooling flue gas below dew point and air-pollution control); 5—the limitations of coal cleaning and boiler size and cost as related to fuel characteristics; 6—future possibilities and conclusions. The development of combustion apparatus for power boilers is progressing at a lively pace. There has been no letup in improvements in design of pulverized-coal-fired boilers, and there is a strong trend at present toward improving dry-bottom units. Spreader stokers with overfire jets and dust collectors as standard equipment are gaining favor. Less than 10 years in commercial use, cyclone burners are going into numerous installations here' and abroad.' Underfeed and traveling-grate stokers have long since been developed for heavy-duty operation, yet new developments in overfire jets and humidification of air blast have improved their performance. A water-cooled vibrating-grate stoker of German origin is being introduced into the United States and Canada." The primary objectives of an ideal coal combustion device are: capacity to burn the variety and sizes of coals likely to be economically available during the life of the unit; capacity to burn the coals automatically for a wide load range and rapid load fluctuations and to burn the coals completely to CO2, H2O, and SO2, which means without smoke and cinders, or carbon in the refuse; capacity to control and discharge all the ash in final granular form without ash adhesion to walls or tubes, and without flue dust; minimum furnace volume; minimum labor and maintenance; low initial and operating cost. Regardless of the method of burning, the gaseous products of coal combustion are N2, CO2, O2, H20, and SO?. By way of illustration, the coal analyses in Table I is assumed from an installation described by E. McCarthy.' When coal is burned with 20 pct excess air (theoretical air, 9.23 lb per lb of coal), the quantities of combustion gas shown in Table II are produced. In addition, the gases carry particles of fly ash, unconsumed cinders, soot particles, and small but significant amounts of vaporized oxides and sulphates of sodium, potassium, lithium, phosghorous, iron, and other metals. In recent years, germanium, one of the rare metals found in coal, has been shown to oxidize and vaporize at combustion temperatures and to be concentrated by reconden-sation at lower temperatures." Pulverized coal and cyclone flames" have peak temperatures of 3000' to 3500°F. Temperatures in fuel beds of large underfeed stokers reach maxima of 3000°F, sufficient to fuse almost any ash and to volatilize some of it. These peak temperatures are above the optimum necessary for rapid combustion, but they hasten heat transfer for ignition as well as boiler heat absorption. Furnace and gas temperatures increase with combustion air preheat. Low excess air has the same effect. Fine coal pulverization and highly turbulent combustion shorten the distance for fuel burnout, increase flame temperature, and speed up heat transfer. Rates of combustion of pulverized coal exceeding 200,000 Btu per cu ft per hr have been demonstrated in atmospheric gas-turbine combusters,

Jan 1, 1955

By Elmer R. Kaiser

G. A. Vissac (Consulting Engineer, Vancouver, B. C.) —Some of the data presented in this paper, in connection with cost studies of washed coals, should be clarified and qualified. Washing a raw coal down to a low ash content may mean higher losses in refuse, but the cost of the operation may be more than compensated by the increased value of the resultant product. However, careful engineering and accurate machinery are required to insure maximum profits. Washing at a low gravity separation demands a very accurate automatic process, chiefly when the small sizes of coal are being handled. Coal consumers today are buying mostly on a Btu basis, cognizant of the fact that low ash contents result in lower installation costs and in savings in handling and transportation charges. The attached Table X gives the actual balance sheets of the various cleaning operations corresponding to various ash contents for the two coals presented in the paper, using the old technique of a two-products separation. or the new technique, now commonly used in Europe, of a three-way separation made possible by new methods and new equipment. Savings in capital costs, handling, and transportation have been calculated on the basis of a delivered cost f.O.b. plant of $10.00 per ton of coal at 16.5 pct ash (actual 1953 average price reported by Consolidated Edison, N. Y.), and of $3.50 at the mine also for 16.5 pct ash coal. In both cases this corresponds to values of 126 and 4.26 per 1 pct of combustible and per ton, respectively, on delivered and on at the mine prices. As an example, the middle western coal at 16.5 pct ash, raw, washed down to 4 pct ash in a two-way process will give 53.75 pct of clean coal at 4 pct ash and 46.25 pct of refuse at 31 pct ash. The value of the losses will be 0.4625x(100—31)x0.042 = $1.34. Add the cost of washing, capital, and operation, which is $0.35. The total cost will then be $1.69. But with a three-ways separation we will have 53.75 pct of first-grade coal at 4 pct ash, 37.25 pct second-

Jan 1, 1955

By H. Macpherson

Durhm has a well earned reputation for supplying some of the finest coking caals in the world. The caals, in general, vary in rank from 301 to 501/2. Durham has traditionally produced foundry coke for the major proportion of the foundries in Great Britah. Durham coals can be described as clean, soft coals, which yielded, with the old hand methods of coal getting, low ash small coal because the extraneous shale was normally harder than the coal and occurred in larger pieces. Under these conditions, with a plentiful supply of cheap labor, it used to be sufficient, at most collieries, to hand clean the large sizes of coal which could then be remixed with the untreated small coal to produce carbonization coals, ready for the coke ovens, at under 5 pct ash. With the introduction of mechanical methods of coal mining, the coal gradually required more cleaning. The first method adopted to resolve this problem was the introduction of pneumatic dry cleaners for the small coal. Although such machines had little effect on the fine coal (say below 1/8 in.), they could clean the intermediate sizes of coal. This, coupled with hand cleaning the larger sizes above 1-1/2or 2 in., resulted in a combined run-of-mine mixture below 6 pct ash, capable of maintaining the quality and reputation of the cokes produced. In more recent years, the intensification of mechanization and power loading, coupled with gradual exhaustion of the cleaner seams, has created the need for a more complete method of coal cleaning. This particularly applies in the fine sizes (say below 1/50 in.) which normally vary, under present day conditions, between 15 and 35 pct ash and are much too dirty to be included in the raw state in a carbonization mixture. This pronounced change has been accelerated because legislation controlling the dust conditions of mine airs for the prevention of pneumoconiosis has resulted in tk~e extensive use of water underground and a consequent increase in moisture content of the run-of-milie output. The presence of damp fine coal decreast the efficiency of prescreening and dry cleaners, so that this type of preparation for low ash coking coals is decreasing, although it is still used satisfactorily for industrial coals in the medium ash range. Table I shows the gradual increase in mechanization, the reduction in manpower, the increased use of explosives per ton of coal extraction, and the increase in the proportion of coal cleaned by mechanical means in Great Britain. Although similar figures are not available for Durham Div. until after the date of nationalization of the coal industry, it is probable that the increase in mechanical cleaning, particularly by dry cleaners, was more marked in the Durham collieries than elsewhere in the country. As dry clealiers were replaced by coal washeries in the Gritish ccal industry, no special attempts were made to recover the slurry, with the result that large outflows of dirty water were allowed, deposits of slurry came in to lagoons and neighboring streams, and a proportion of fine material was lost from the coking coal. This position, coupled with the higher moisture of the washed coking coal, resulted in adverse effects on coke oven throughputs and coke quality. It is now realized that the natural coal fines are an essential ingredient of coking coals in obtaining the correct coke structure in metallurgical cokes. This, together with economic pressure, led to the introduction of flocculation and filtration plants for the recovery of slurries, and later, when the ash contelits of the filter cakes were too high, to the introduction of froth flotation equipment. After this position had been reached, the tailings from the froth flotation pIants were, in many cases, still allowed to constitute an undesirable effluent. Recent legislation on river pollution has changed this picture; it is now necessary to provide a circuit which is completely closed so far as solids are concerned. The gradual increase in the coal cleaned by wet methods and froth flotation in Durham Div, is shown in Table 11. It is now an accepted feature of new Washeries that. froth flotation should be an integral part of the washing process from the initial installation of the plant.

Jan 1, 1962

By Shiou-Chuan Sun

THIS paper presents the design and operation of a frothmeter capable of measuring the frothing characteristics of pine oils and other frothing reagents. The experimental data show that the froth-ability of pine oil is governed by: 1—rate of aeration, 2—time of aeration, 3—height of liquid column, 4—chemical composition of pine oil, 5—pH value of solution, 6—temperature of solution, and 7—concentration of pine oil in solution. The effect of mineral particles on the behavior of froth also was studied, and the results can be found in a separate paper.' The results also show that the relative froth-abilities of pine oils in the frothmeter generally correlate with those in actual flotation, provided that other factors are kept constant. In addition to pine oils, the other well-established flotation frothers were tested, and the results are included. In this paper, compressed air frothing is the frothing process performed by means of purified compressed air, whereas sucked air frothing is the frothing process accomplished by purified air sucked into the glass cylinder by a vacuum system. The term vacuum frothing denotes that froth was formed by degassing of the air-saturated liquid under a closed vacuum system. Apparatus The frothmeter, shown in Fig. 1, is capable of re-producibly measuring the volume and persistence of froth as well as the volume of air bubbles entrapped in the liquid and is capable of being used for compressed air frothing, sucked air frothing, and vacuum frothing. Fig. la shows that for compressed air frothing, the apparatus consists of an airflow regulating system, 1-3; a purifying and drying system, 4-8; a standardized flowmeter to measure the rate of airflow from zero to 500 cc per sec, 9; and a graduated glass cylinder, 13; equipped with an air regulating stopcock, 10; an air chamber, 11; and a fritted glass disk to produce froth, 12. The fritted glass disk, 5 cm in diam and 0.3 cm thick, has an average pore diameter of 85 to 145 microns. The pyrex glass cylinder has a uniform ID of 5.588 cm and an effective height of 63 cm. The inside cross-sectional area of the glass cylinder was calculated to be 24.53 sq cm, or 3.8 sq in. For sucked air frothing, Fig. lb shows that the apparatus for compressed air frothing is used again, with the following modifications: 1—compressed air and its regulating system, 1-3, are eliminated; and 2—a vacuum system, 16, equipped with a vapor trap, 15, and a vacuum manometer, 17, is added. The vacuum system can be either a water aspirator or a laboratory vacuum pump. Any desired rate of airflow can be drawn into the glass cylinder, 13, by adjusting the opening of the air regulating stopcock, 10. The sucked air stream is cleaned by the purifying and drying system, 4-8, before entering the glass cylinder, 13. When this setup is used for vacuum frothing, the air regulating stopcock is closed. The frothmeter has been used for almost 3 years and has proved to give reproducible results, as illustrated in Table I. With a magnifying glass and suitable illumination, the frothmeter also can be used to study the attachment of air bubbles to coarse mineral particles.' Experimental Procedures Except where otherwise stated, the data presented were established by means of the compressed air method. The volume and persistence of froth were recorded respectively at the end of 4 and 6 min of aeration at a constant rate of airflow of 29.3 cc per sec, which is equivalent to 71.6 cc per sq cm per min, or 462.6 cc per sq in. per min. The aqueous solution for each test, containing 1000 cc of distilled water and 19.2 ± 0.5 mg frothing reagent, was adjusted to a pH of 6.9 0.2. The volume of froth is expressed as cubic centimeter per square centimeter and is equivalent to the height of the froth column (the distance between the bottom and the meniscus of the froth). The volume of froth was obtained by multiplying the height of froth by the cross-sectional area of the glass cylinder, 24.53 sq cm. Before each test, the glass cylinder, 13, was cleaned thoroughly with jets of tap water, ethyl alcohol, tap water, cleaning solution, tap water, and finally distilled water. The cylinder with stopcock,

Jan 1, 1953

By Shiou-Chuan Sun

THIS paper presents the design and operation of a frothmeter capable of measuring the frothing characteristics of pine oils and other frothing reagents. The experimental data show that the froth-ability of pine oil is governed by: 1—rate of aeration, 2—time of aeration, 3—height of liquid column, 4—chemical composition of pine oil, 5—pH value of solution, 6—temperature of solution, and 7—concentration of pine oil in solution. The effect of mineral particles on the behavior of froth also was studied, and the results can be found in a separate paper.' The results also show that the relative froth-abilities of pine oils in the frothmeter generally correlate with those in actual flotation, provided that other factors are kept constant. In addition to pine oils, the other well-established flotation frothers were tested, and the results are included. In this paper, compressed air frothing is the frothing process performed by means of purified compressed air, whereas sucked air frothing is the frothing process accomplished by purified air sucked into the glass cylinder by a vacuum system. The term vacuum frothing denotes that froth was formed by degassing of the air-saturated liquid under a closed vacuum system. Apparatus The frothmeter, shown in Fig. 1, is capable of re-producibly measuring the volume and persistence of froth as well as the volume of air bubbles entrapped in the liquid and is capable of being used for compressed air frothing, sucked air frothing, and vacuum frothing. Fig. la shows that for compressed air frothing, the apparatus consists of an airflow regulating system, 1-3; a purifying and drying system, 4-8; a standardized flowmeter to measure the rate of airflow from zero to 500 cc per sec, 9; and a graduated glass cylinder, 13; equipped with an air regulating stopcock, 10; an air chamber, 11; and a fritted glass disk to produce froth, 12. The fritted glass disk, 5 cm in diam and 0.3 cm thick, has an average pore diameter of 85 to 145 microns. The pyrex glass cylinder has a uniform ID of 5.588 cm and an effective height of 63 cm. The inside cross-sectional area of the glass cylinder was calculated to be 24.53 sq cm, or 3.8 sq in. For sucked air frothing, Fig. lb shows that the apparatus for compressed air frothing is used again, with the following modifications: 1—compressed air and its regulating system, 1-3, are eliminated; and 2—a vacuum system, 16, equipped with a vapor trap, 15, and a vacuum manometer, 17, is added. The vacuum system can be either a water aspirator or a laboratory vacuum pump. Any desired rate of airflow can be drawn into the glass cylinder, 13, by adjusting the opening of the air regulating stopcock, 10. The sucked air stream is cleaned by the purifying and drying system, 4-8, before entering the glass cylinder, 13. When this setup is used for vacuum frothing, the air regulating stopcock is closed. The frothmeter has been used for almost 3 years and has proved to give reproducible results, as illustrated in Table I. With a magnifying glass and suitable illumination, the frothmeter also can be used to study the attachment of air bubbles to coarse mineral particles.' Experimental Procedures Except where otherwise stated, the data presented were established by means of the compressed air method. The volume and persistence of froth were recorded respectively at the end of 4 and 6 min of aeration at a constant rate of airflow of 29.3 cc per sec, which is equivalent to 71.6 cc per sq cm per min, or 462.6 cc per sq in. per min. The aqueous solution for each test, containing 1000 cc of distilled water and 19.2 ± 0.5 mg frothing reagent, was adjusted to a pH of 6.9 0.2. The volume of froth is expressed as cubic centimeter per square centimeter and is equivalent to the height of the froth column (the distance between the bottom and the meniscus of the froth). The volume of froth was obtained by multiplying the height of froth by the cross-sectional area of the glass cylinder, 24.53 sq cm. Before each test, the glass cylinder, 13, was cleaned thoroughly with jets of tap water, ethyl alcohol, tap water, cleaning solution, tap water, and finally distilled water. The cylinder with stopcock,

Jan 1, 1953

By Shiou-Chuan Sun

DISCUSSION shiou-chuan sun (author's correction) - Too late for revision, the author realized that the data on stability of froth presented in this paper were not based on the same height of froth column. It is recommended herein that the stability of froth is better calculated by dividing the time elapsed until the froth disappeared by the volume (in cu cm per sq cm) or height of the froth column, and is expressed in sec per cm. The stability data in Tables I, II, and III of this paper were recalculated, in accordance with this recommendation, and the results are shown respectively in the accompanying Tables VII, VIII, and IX. Compared with Tables I, II, and III of the paper, the stability data of Tables VII, VIII, and IX are more reliable in disclosing the persistence of froth of various frothers having different frothabilities. For example, the conclusion given in the paper that the stability of froth is generally proportional to the volume of froth becomes invalid when the stability of froth is expressed in sec per cm, as shown in Tables VII, VIII, and IX. Table VII shows that both the stability and volume of froth increase with the addition of small amounts of n-hexyl alcohol, and that further additions of n-hexyl alcohol decrease the stability but increase the volume of froth. Table VIII indicates that the stability of different types of frothers is governed by their chemical composition and is difficult to predict from their frothability. Table IX shows that the stability of different frothers with similar nonpolar groups tends to increase with a decrease of the hydro- philic property of their polar groups, and consequently with the decrease of their solubilities. This is further substantiated by the fact that the stability of frothers, as listed in Table IX, is generally lowered by the solvent effect of absolute ethyl alcohol. Table IX shows also that the stability of froth tends to increase with the decrease of the volume of froth* provided the frothers used have similar nonpolar groups and dissimilar polar groups.

Jan 1, 1953

By Shiou-Chuan Sun

DISCUSSION shiou-chuan sun (author's correction) - Too late for revision, the author realized that the data on stability of froth presented in this paper were not based on the same height of froth column. It is recommended herein that the stability of froth is better calculated by dividing the time elapsed until the froth disappeared by the volume (in cu cm per sq cm) or height of the froth column, and is expressed in sec per cm. The stability data in Tables I, II, and III of this paper were recalculated, in accordance with this recommendation, and the results are shown respectively in the accompanying Tables VII, VIII, and IX. Compared with Tables I, II, and III of the paper, the stability data of Tables VII, VIII, and IX are more reliable in disclosing the persistence of froth of various frothers having different frothabilities. For example, the conclusion given in the paper that the stability of froth is generally proportional to the volume of froth becomes invalid when the stability of froth is expressed in sec per cm, as shown in Tables VII, VIII, and IX. Table VII shows that both the stability and volume of froth increase with the addition of small amounts of n-hexyl alcohol, and that further additions of n-hexyl alcohol decrease the stability but increase the volume of froth. Table VIII indicates that the stability of different types of frothers is governed by their chemical composition and is difficult to predict from their frothability. Table IX shows that the stability of different frothers with similar nonpolar groups tends to increase with a decrease of the hydro- philic property of their polar groups, and consequently with the decrease of their solubilities. This is further substantiated by the fact that the stability of frothers, as listed in Table IX, is generally lowered by the solvent effect of absolute ethyl alcohol. Table IX shows also that the stability of froth tends to increase with the decrease of the volume of froth* provided the frothers used have similar nonpolar groups and dissimilar polar groups.

Jan 1, 1953

By R. H. Jamison

A relatively new haulage system is described. Employed by the Delmant Fuel Co.. the "Full Dimension" system provides an uninterrupted flow of coal from a loader or continuous miner at the face to the main line transportation system. This system is said to provide a higher percentage of recovery as well as additional safety and production. Delmont Fuel Co. is employing a comparatively new system of transportation known as a Full Dimension system. Cne of these systems has been in operation for a year at the company's 10-B Mine as a part of a conventional section. A second was installed at the No. 10 mine in late 1960 to handle the production of a Colmol in a pillar section. SYSTEM COMPONENTS A Full Dimension system is a haulage system that provides an uninterrupted flow of coal from a loader or continuous miner at the face to the main line transportation system. The equipment required for this system consists of a series of interconnected chain conveyors that are mobile and articulated. They will retract or extend a sufficient distance for the development of a five-entry system; or, in the Colmol pillar section, it provides reach of 210 ft in all directions from the section belt. The components of this system are: l)One 160-ft chain line placed in tandem with the belt conveyor. It has a self-propelled drive, is 20 in. wide and 9 in. deep. Moving this conveyor requires the assistance of a loading machine or cutting machine. 2) One 40-ft piggyback that discharges along the entire length of the 160 ft chain conveyor. 3) A mobile bridge carrier, which is a self-propelled conveyor with four wheel steer and four wheel drive, twenty-eight feet long, it delivers coal to the receiving end of the piggyback. Axles steer individually making possible almost lateral movement. 4) Another 40-ft piggyback, duplicate of item 2 that delivers coal along the entire length of item 3 (mobile bridge carrier). 5) A second mobile bridge carrier, similar to the first, which deliver coal to the piggyback (item 4). 6) A third 40-ft piggyback, duplicate of items 2 and 4. This pig is attached to the loading machine and delivers its coal along the length of the second mobile bridge conveyor. Since the original preparation of this paper, the Delmot Fuel Co. has been able to eliminate the 160-ft chain conveyor. This was accomplished by connecting the outby piggyback directly to a loading machine with an extended boom. The loading machine loads directly onto the belt. This change has resulted in a substantial reduction in moving time and greatly increased flexability. A single trailing cable powers the entire string of equipment. It is attached to the side of the equipment in such a way as to keep it off the ground and afford maximum protection. The tramming rate of this equipment is 90 fpm. The conveyor capacity in a conventional section at Delmont's mines is 7.5 tpm and in the Colmol section is 5.5 tpm. This regulation is a simple function of conveyor speed. To visualize operation of this equipment, it would be well for me to touch briefly on local conditions in the Upper Freeport seam in which we mine. (Also, see the photographs of some of the equipment in use.) DELMONT'S TOPOGRAPHY The Delmont Fuel Co. operates two mines in this seam in Westmoreland County, Pa. The No. 10 mine, which was opened in about 1912, is now almost worked out. Depending on economics in the industry, it has a life of two to four years on a declining production basis. A year ago a new drift mine was opened which is called No. 10-B. It is about two miles from the cleaning plant and is connected thereto by an overland belt conveyor. The new mine is being developed at a rate calculated to take up the slack as the old mine plays out. The Upper Freeport seam averages 4.2 ft in thickness in the area of the Delmont mines. It carries 4 in. of boney coal at the top of the seam and a middle man of from 2 to 4 in. We mine just above a 1-in. slate parting which has 4 to 6 in. of highly laminated coal beneath it. This material normally makes a very firm bottom. The roof varies from dark shale to sand rock and 36-in. bolts are placed on 4-ft centers for roof support. All working places are driven 20 ft wide on development and 25 ft wide on retreat. Selection of mobile chain conveyor equipment when it became available, was a very natural move for Delmont Fuel to make, because chain conveyors and piggybacks had been in use at the company's mines for about 12 years. Grades in the new mine

Jan 1, 1961

By H. B. Charmbury, D. R. Mitchell

Recovery of fine coal from solids reporting to wash water has become increasingly important. These solids range from about 28 mesh to 0 and constitute 3 to 7 pct by weight of the feed tonnage to a coal preparation plant. In the past they were usually wasted to streams, but present clean stream legislation makes it necessary to remove them from the wash water and, if economically possible, recover the associated coal as a saleable product. Where these solids are not wasted to streams or to slime ponds they are recovered by thickeners and filters. Since filter cake is often high in ash or sulfur, or both, it is not readily marketed or used. Common practice is to mix it with other plant products, generally slack or cleaned run-of-mine. Uniform blending is difficult, not only because the impurity is usually high, but also because during certain operating periods it is possible to get a concentration of this filter cake in a railroad or barge. At some plants filter cake has been derived from the froth flotation clean-up of Dorr thickener underflow. The operating circuits required for a froth flotation plant make this a comparatively high-cost treatment, and in the last few years experiments have been carried out with bulk oil methods. These also are quite expensive and lend themselves only to special situations. Table I gives some characteristics of thickener underflows for plants in the Pittsburgh area treating Pittsburgh seam coal. It will be noted that there is a concentration of impurities in the -200 mesh sizes.

Jan 1, 1960

By A. Brown

RAPIDLY rising world demand for mineral products has accelerated depletion of the more readily accessible ores, particularly those of premium grade. Operations must proceed at a faster rate to deeper horizons, where there are problems of ground control that restrict economic recovery of the mineral and introduce new hazards to mine personnel. For some of the new Canadian mining fields severe problems of ground stress are still in the future, but much of Canada's mineral output comes from well established mines that have reached considerable depths. It must also be anticipated that the depletion of premium grade ores will eventually necessitate mining low grade material at competitive cost, and unsolved problems of ground control can materially increase costs. Studies of ground stress phenomena in underground workings are therefore considered a present-day necessity, especially since these investigations must be a long-term effort owing to the highly complex nature of the problem and the difficulty of making full-scale observations. In recent years the Mines Branch of the Department of Mines and Technical Surveys in Canada has conducted such an investigation. This report of the work is primarily concerned with stress problems in some Canadian coal mines and, in particular, with mine bumps. Mines Branch investigations of this and related problems are carried out by a small team of mining engineers and physicists who, with assistance from geologists, pool their special skills in an integrated program. These Canadian investigations combine a field and laboratory approach. Since data obtained under contrasting mining conditions allow useful comparisons of ground behavior under stress, this article will present data gathered under a wide range of Canadian coal mining conditions. Two contrasting longwalls in eastern Canada will be discussed and two dissimilar pillar extraction operations in western Canada. Locale of Observations In eastern Canada observations are being made in the collieries at Springhill, Nova Scotia. The district's No. 2 mine, which has a history of bump occurrences dating back to 1917, is currently being worked at a depth of 4200 ft. The mining method is longwall retreating along the strike and the practice is to retreat three walls in line, each about 350 ft long. There is no machine cutting or shooting—the coal is hand-picked and hand-loaded onto face conveyors. There are three rows of hardwood chock block supports parallel to the coal face. These supports are built during the two daily coal loading shifts, and the rear line is withdrawn during the midnight shift. The goaves are partially packed by stone midwalls 12 ft wide which are built forward daily in step with the face advance. The seam dips at 15° and averages 9 ft of strong, well cleated bituminous coal enclosed in competent beds of shales, sandy shales, and sandstones of Carboniferous age. The bumps rarely disturb the retreating longwall faces but manifest themselves as violent upthrusts of pavement in the levels leading to the walls. Observations have also been made in Springhill's No. 6 mine, which operated in a 5-ft seam of coal at a depth of about 2500 ft. Although No. 6 mine was

Jan 1, 1959

By E. C. Olsen

This paper describes the unusual hazardous conditions encountered in mining thick, inclined coal beds under heavy cover in Carbon County, Utah. This description includes heavy roof conditions, sloughing ribs, steep pitches, and severe bouncing experience. Emphasis is placed on the top bench system of mining and development in the top half of a 16-ft seam, under the roof rock, which is bolted, and the recovery of the bottom coal on retreat; success of bolted roof and ribs; and shuttle car haulage on pitches up to 25%. The use of protective cages for operators of face equipment, the guarding of belt conveyor systems to prevent injury associated with deliberate entry, and the positive lockout of electrical equipment is described and illustrated. Most coal mining areas of the Western United States are characterized either by thick beds, steep pitches or heavy cover. Individually, each of these may present inherent safety hazards that in- fluence mining. When all are met at one location, the combined effect is to make the safe operation of the mining property a challenge to management. These combined conditions are met in the coal mines of the Columbia-Geneva Steel Div. of the United States Steel Corp., in southeastern Utah. This coal bed measures up to 16 ft in thickness, the pitch is as much as 25% and mining is under a cover approaching 3000 ft. The Sunnyside coal bed is found in the Book Cliff Mountains, a great escarpment extending from west of Castle Gate, Utah, some 180 miles to Grand Junction, Colo. West of Green River, Utah, the Book Cliffs swing sharply to the north because of the domelike uplift of the San Rafael Swell to the southwest. It is in this northerly extension of the Book Cliffs that the Sunnyside coal in the Blackhawk Formation of the Mesa Verde Group, Upper Cretaceous Age is found. The strata are transected by numerous faults with from a few inches to as much as a hundred feet of displacement. All of these faults have had some effect on the mining operation, either in roof control or, in the case of major faulting, the pattern of mine development. MINING PRACTICES Mining has generally been confined to the Lower Sunnyside Seam which measures from 7 to 16 ft in

Jan 1, 1963

By M. H. Shumate

Rope-type haulage has had many applications in the mining and allied industries. Records have indicated favorable results both from a standpoint of efficiency and investment. The Truax-Traer Coal Co. has used some variations of rope haulage at their mines and preparation plants. They have built a system which solves their particular problems, and have an investment which has indicated an annual savings in operating costs. Rope haulage, as applied to the mining industry, goes back many years for both underground and surface mining, or a combination of both. The Truax-Traer Coal Co. has used gravity retarding hoists as well as several variations of rope haulage, including electrical and combinations of both, but each was limited in its use and application by natural conditions and economies of the operation. Several systems of rope haulage equipment are offered by manufacturers today for handling railroad car movement on a limited or continuous basis. The portal and railroad loading facilities of Truax-Traer Coal Co.'s Burning Star Slope mine, located in Jackson County, Ill., were moved in 1960. The old location was abondoned, eliminating 3% miles of underground track haulage. The mine was converted to an all-belt system, and coal is loaded at a new raw coal plant and shipped by railroad cars to a central cleaning plant. The company wanted to operate the surface facilities as efficiently as possible, employing a minimum number of workers, using the latest type railroad car movers. The existing rope haulage facilities at several locations throughout the country were examined and considered for the Slope mine location. The application of each appeared favorable but lacked flexibility, and it was difficult to justify the capital investment. The company decided to investigate the possibility of building a system that would apply to their problem, and have an investment that could be amortized in a relatively short period. SPECIFICATIONS They employed the services of Allen and Garcia Co. of Chicago, Ill. Through combined efforts, a railroad car rope system was designed to specifications as shown in Fig. 1. The Falk Corp. of Milwaukee, Wis. built the hoist to the specifications as shown in Fig. 2. It has an all welded base and pedestals. The all welded drum is not grooved. A single helical gear was used in lieu of herring bone type. A Falk speed reducer, Unit 90Y3-A, is driven by a 25-hp 1800-rpm, 440-V ac type 'C' (high starting torque) drip-proof, Frame 324-U motor. The speed reducer shaft is equipped with a solenoid operated ac spring set shoe brake, operating on a 7-in. diam. brake wheel. The dolly car, shown in Fig. 3, was field constructed, using the trucks and frame of a railroad tank car. Truax-Traer did not alter the frame but added plates and anchors for hoist ropes and frames for 8.5 tons of concrete to be used as ballast on the car. The limit switch operating shoe was also installed and the car coupler latch mechanism overhauled. The hoist rope selected was 1% in. in diam., 6x37 improved plow steel extra flexible, right lay, regular lay, with independent wire rope core, and preformed. Wedge sockets were used to connect rope to the dolly car as shown in Fig. 4. The system employs two 30-in. and three 16-in. diam. sheaves. All are bronze bushed and equipped with grease fittings. Larger sheaves are used for directional change of hoist rope and the smaller ones as snub sheaves to correct fleet angle of rope as it approaches the hoist drum. The Trench lay cable, buried between the two tracks, is used for electrical distribution and controls between hoists and operator's cab. Foundations for the system required 187 cu yards of reinforced concrete using approximately 60 cu yd each at three locations. Four hoists are employed in the system, two for each track. A pair is interconnected electrically and when one operates, moving the cars, the tail hoist idles with brake released. The process is repeated for the reverse direction. A limit switch, mounted in the track near each end of the system, is tripped by the dolly car as it approaches the hoist. The limit switch controls movement in one direction only, protecting the dolly car and establishing positive con-

Jan 1, 1962