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By Shiou-Chuan Sun

THE effects of oxidation on the flotation behavior of sulphide minerals have been extensively studied 1-3 but no similar study has been made of coals. Coals of bituminous and lower rank undergo atmospheric oxidation during mining and storage, and their degree of oxidizing increases with temperature, 4,5 and time of exposure." The problem of recovering fine oxidized coal from culm piles, rivers, and streams by froth flotation has become increasingly important to the coal industry as well as to those concerned with air and stream pollution. This paper describes the flotation behavior of oxidized and unoxidized coals and carbons in the presence of oily collectors and presents the reagents found effective in the flotation of oxidized coals and carbons. Except where otherwise stated, flotation tests were performed in a 50-g Denver flotation cell. Solid feed for each test was 30 g, and the pulp dilution was kept at 8.5. The —65 mesh coal sample was wetted by agitation with water in the flotation cell for 6 min, then conditioned with a predetermined amount of collector for 2 min, and finally floated with pine oil as frother for 3 min. The oxidation of coals and carbons was performed in a vertical, electrically heated glass tube, shown in Fig. 1. One hundred grams of —65 mesh coal were fed into the glass tube and heated to 10° to 20 °C below the predetermined temperature of oxidation. Compressed air was admitted to the tube at a rate of 0.05 cu ft per min and the temperature was controlled by a variable voltage regulator. The oxi-

Jan 1, 1955

[Abstmct of remarks by R. II. Sweetser arid suhscqueut discussion at the New York Meeting, February, 1926. The Transactions mill ultimately contain a report of work resulting from these discussions. Need for a standard of ualuration.—There should be an accepted method for evaluating the different kinds of bituminous coals based on their carbon just as that for iron ores is based on iron. Variation in the classes of coals and their desirability fol the different industrial uses is analagous to the case of iron ores. The iron people have developed a standard. In the past coals have not been intelligently distributed because the public has not been informed as to its uses. Many operators have known nothing about the final disposal of the coal shipped from their mines. Sales agents and retailers have had no means of evaluating coals except to make the price as high as the market will stand. Producers and distributors need to know how much more valuable the coals would be if prepared so as to contain less objectionable material. Mr. Swectser cnlletl on the Goal and Coke Committee of the Institnte to organize a study through producers and users to develop a method scientifically and comnereially sound E. A holbrook said that the information exists but the public is unwilling to investigate. He praised Ashley's1 nomenclature of the different kind of coal which has been accepted by the Coal Mining Institute of America and recommended to the American Enginecring Standards Commitee. G. S. Rice pointed out the wider variety and distribution of consumption of coal compared to Iron ore. C'. M. Young approved Mr. Sweetser's remarks and hoped lor a research institution, probably endowed, to study utilization of coal after it comes from the ground. Mr. Sweetser cited incidents to prove his former statements. He said that a reduction of 1 per cent. ash in the coal would cut down the cost of pig iron at a blast furnace 30 c. a ton, that the coal men should tell what the iron men would have to pay to get better coal. H. J. Prryn endorsed the need of blast-furnace operators to know the kind of coal from which their- coke is made and the latler's ash content, as this affects coke consumption. A. C. Fieldncr explained that the difficulty of evaluating coal on the basis of chermical analyses for calorimetric tests is: (1) representativr

Jan 1, 1927

By F. J. Chernosky

Most previous studies of coal flotation utilized chemically pure reagents. Since such reagents are not available in quantity, a study of various wmmercially available reagents as frothers was undertaken. Flotation procedures similar to those used in a large number of commercial coal flotation plants were employed. The reagents used were compared on yield-selectivity-wst basis. The selectivity was found to be relatively constant with yields above 70% but the reagent cost varied greatly. The types of reagents available were compared. In the past, a number of papers have been written on the use of various reagents as coal flotation frothers. For the most part, these have been an attempt to correlate the intrinsic chemical and physical proper- ties of the reagents or reagent groups to the floata-bility of various fine coals. Because of the type or nature of evaluations made, chemically pure reagents were normally used. Generally, these reagents are unavailable in commercial quantities. Therefore, the results obtained with these reagents were usually limited in usefulness to the plant operator or metallurgist. This paper presents the results of a study in which commercially available forms of various reagents were used as flotation frothers. Flotation procedures which would conform in general to those employed in a large number of commercial coal flotation plants were utilized. The objective of the test was to compare these reagents on yield-selectivity-cost basis, that is, 1) the yields obtained, 2) the ash levels obtained and 3) the reagent cost relating to these results. REAGENTS The reagents employed have been grouped into the following chemical classes: 1) aliphatic alcohols, 2) terpines, 3) glycols (synthetic water-soluble

Jan 1, 1963

By S. A. Braley

DRAINAGE water from coal mines is probably the most serious water pollution problem today, varying in importance according to location of the mines and geological structure. Drainage may be either acid or alkaline in character. Acid discharge, the most severely detrimental to a stream, is caused by natural oxidation of the sulfuritic material (FeS2) in the strata associated with the coal seam. Since the acid is the result of a natural reaction the acid water differs because it does not cease with abandonment of the mining operations. There is no known economical method of neutralizing acid mine water or any practical method to prevent oxidation of exposed pyrite. Since production of acid from a mine does not stop when mining stops, the total quantity produced depends entirely upon the excavated areas. The increasing volume of acid water in manv mines has greatly increased operating costs. Pumping is expensive and acid mine waters are destructive of all equipment, especially metals in pumps and piping, and necessitate the use of corrosion-resistant materials. Discharge of the acid mine drainage into streams neutralizes their normal alkalinity, causes them to become acid, and produces an environment unfavorable for aquatic life and unsuited for industrial or domestic use without costly treatment. Mine drainages vary in percentage composition over wide limits, although the usual dissolved substances are ferrous and ferric iron, aluminum, calcium and magnesium sulfates, and lesser amounts of sodium, potassium and manganese sulfates and chlorides. Some alkaline discharges may contain heavy concentrations of iron as iron bicarbonate. These waters may produce iron hydroxide deposits in the receiving stream but do not cause it to become acid. In the extensive literature on acid mine water there appears to be a great deal of confusion about the importance of various components and the methods for their determination. In many instances faulty conclusions have been drawn from use of unsuitable methods of analysis. It is desirable that the factors and terms used in evaluation of analyses of mine waters should be so clearly defined that any interested person could properly appraise any analytical report. Some analysts report complete chemical analyses of mine waters and neglect to record drainage volumes. Others report only a minimum of analytical data after taking no precautions to preserve the original composition of the water during the time elapsing between collection and analysis. Some use methods of analysis intended for so-called pure waters of the potable and boiler water classes. These methods are not applicable to highly buffered waters such as mine water. Probably the most common criteria for evaluation are pH, free acidity, or acidity or alkalinity to methyl orange or methyl red, total acidity or acidity to hot phenol-phthalein, and the sulfate content. If these determinations are made on carelessly collected samples after a few days to weeks standing in warm rooms, they do not in any way represent the character of the water flowing from the mines. It is hoped that a brief discussion of the fundamental value of some of these factors may lead to a bettqr understanding of the need for more careful evaluation of mine water discharges. The term pH is one used by chemists to express relative acidity or alkalinity in terms of concentration of effective hydrogen ion in a solution. It is defined as the negative logarithm of the hydrogen ion concentration or activity in equivalents per liter. pH = logarithm A neutral solution, which is one containing the same number of hydrogen and hydroxyl ions, has a pH of 7. As the hydrogen ions increase and the solution becomes more acid, the pH decreases toward zero; and as the hydroxyl ions increase and the solution becomes more alkaline, the pH approaches 14. When dissolved in water to a dilute solution acids like sulfuric and hydrochloric, commonly known as strong acids, ionize completely, and the pH or hydrogen ion concentration varies with molar concentration of the dissolved acid. However, in high concentrations of such acids, the pH or hydrogen ion concentration is less than the acid concentration because the acid does not completely ionize. In only very dilute solutions does the pH represent the total amount of acid that can be neutralized by an alkali. All ionization reactions are equilibrium reactions. If other chemicals are added to the solution of an acid and the added chemical produces an ion that is the same as one of the ions of the acid, the degree of ionization of the acid is altered and the pH changes to some value that represents the active hydrogen ion of the new solution. Thus if iron sulfate is added to a solution of sulfuric acid the pH increases, since the common sulfate ion suppresses the degree of ionization of the sulfuric acid and decreases the effective hydrogen ion. However, the total acidity of the solution is increased. There are two salts composed of iron and sulfate— ferrous and ferric sulfate. In these salts there is no hydrogen ion that can ionize to give an acid solution, but when they are dissolved in water, the pH is less than 7 and the solution becomes acid. This is caused by a reaction known as hydrolysis and is represented by the equations FeSO4 + 2H2O ? Fe(OH)2 + 2H + SO,: or Fe,(SO,)3 + 6H2O ? 2Fe(OH)3 + 6H+ + 3504 A solution with a total acidity of 5000 ppm according to the first equation will have a pH of 4.40 but one with an acidity of 75 ppm, according to the sec-

Jan 1, 1958

By L. Valentik

Many attempts have been made during the last 40 years to evaluate the performance of gravity separation equipement, that is, the effectiveness with which light and heavy particles are separated. The most comprehensive treatment of the subject was made by Cerchar at the 1st International Conference on Coal Preparation held in Paris in 1950. The methods suggested by the Conference were accepted and very widely used in the last two decades. This paper discusses an improved method of evaluation in the light of the now-accepted standard presentation. The float-and-sink analysis of the product is presented on a Gaussian distribution curve, resulting in an easier visualization of the inherent difficulties of separation. The ogives of the distribution curve me then plotted, giving a quantitative measure of the deviation from perfect separation as an error distance instead of an error area. Illustrations of the new method are given both for gravel and for coal preparation, but the content is valid and applicable to other types of minerals which are separated by gravity methods. Many attempts had been made during the last forty years to evaluate the performance of heavy-media separation (HMS) equipment, that is, the effectiveness with which floats and sinks are separated.'-' The most comprehensive treatment of the subject was made by Cerchar at the 1st International Conference on Coal Preparation held in Paris. 6 The primary aim was the thorough understanding of the mechanism of separation and the unified presentation of data on gravity separation so that the evaluation and comparison of washery performance could be made from all over the world. No strict overall standardization has been achieved, but after the conference a more or less uniform presentation of performance was accepted, which, during the last two decades, has been very widely used. In this paper, illustration of the old methods and an improved method of evaluation will be given. HEAVY-MEDIA SEPARATION (HMS) PERFORMANCE CRITERIA In the ideal HMS process, all material lower in density than the specific gravity of separation (SGS) would be recovered as floats and all material of higher density would appear as sinks. In order to evaluate the misplaced material, the washery products are tested at the density at which the washing unit is operated. The original type of plot1,7, 8 is shown in Fig 1; this was developed primarily for coal cleaning units. The curve for raw coal represents the cumulative percentages of sink material. The refuse curve is also plotted as a cumulative sink, the percentages being expressed in terms of raw coal. This diagrammatic representation of the results of washing units has the merit of easy visual observance of the degree of separation obtained. The error areas (cross-hatched) are a measure of the amount of misplaced material and therefore they can be used to characterize the quality of separation. The ideal and actual separating performance between floats and sinks can be best seen from the partition curve developed by Tromp,2 where the ordinate is the percentage recovery of the sinks, and the abscissa is the specific gravity (Fig. 2). It can be seen from the shape of the curve that as the SGS is approached, the proportion of material reporting to the improper product increases rapidly. In fact, the SGS can be defined as the density of the material in the feed that is distributed equally between float-and-sink products. When the upper half of the curve is inverted, a shape similar to that of a Gaussian error distribution curve is obtained and therefore the analysis of gravity separation may be carried out by using the law of probability. The shape of the curve in Fig. 2 is determined partly by the density composition of the feed, and partly by the sharpness with which the unit separates floats from the sinks.9, l0

Jan 1, 1970

By T. C. Shelton, W. M. Huang

One of the problems arising from the use of modern machines and systems in coal mining is increased emission of explosive gases into the mine passageways. Interest in degasification of coal seams has grown. The authors summarize the results of coal permeability tests done at Virginia Polytechnic Inst. on one of the fundamental aspects of the problem. There are undoubtedly many factors that affect the emission of gas from coal but the permeability of coal to gas flow would seem to be fundamental. However, there is very little information on coal permeability to gas flow available in the U.S. at the present time, despite the fact that such knowledge would appear to be significant in any degasification effort. Studies made recently in the mining engineering laboratories at Virginia Polytechnic Inst. have determined values of permeability on 1-in. cubes of selected coals. COAL PERMEABILITY The permeability values given in this paper are based on adaptations of Darcy's law formulated by Henry Darcy in I856 from experiments with fluid flow in porous media.' Using air as the gaseous fluid, they were obtained by employing similar methods as are used to dete1:mine the permeability of petroleum and gas reservoir rock specimens. Permeability is. a measure of the ease with which fluids flow through a porous substance, and it is governed by the connected pores within a porous medium. If a few large-sized pores, fractures or channels are present within the medium, the medium can have great permcnability without significant porosity. It is reasonable to assume, therefore, that the permeability of coal, due to its structure, is based on fractures connecting the pore space left as the result of some degree of devolatilization or other changes of the substance from which coal was formed. The results of this irvestigation tend to validate this assumption in that the values obtained indicate a close relationship between the rank and permeability of coal. COALS TESTED AND PROCEDURE The coals whose permeability values were determined in this investigation are given in Table I.

Jan 1, 1962

By J. P. Newell, R. W Storey

This paper describes the development of a continuous float-and-sink process to produce coal low enough in ash content to be suitable for production of electrodes. The cleaned coal had a combined iron and silicon content of 0.239 pct. The efficiency of recovery was 84 pct. AUGER mining is a form of continuous mining in that it completely replaces with a one-cycle operation the older conventional cut, drill, shoot, and load method of mining. Relatively new, having been utilized only within the last few years, it is even more continuous in its production of coal than some other forms of continuous mining being practiced in this country today. There are at present coal recovery augers in operation working along strip-mined highwalls from the outside. This paper will deal with underground mining, specifically with an experiment which was carried out with a machine developed for this purpose. The location of this experiment was in Pike County in eastern Kentucky in the upper split of the No. 3 Elkhorn seam. Coal height was 32 in. and roof conditions were very bad, with 10 in. of soft draw slate coming down with every cut in conventional mining. The coal is high-grade bituminous, excellent for metallurgical coking. Underground auger mining fills a need for a method of recovering coal in thin seams, under bad roof conditions or in split seams. In many cases it might save money by eliminating the handling of excessive quantities of rock. In fact, if the coal can be recovered, as in the case of this experiment, without mixing with any impurities outside the seam, it may be possible to load or ship a satisfactory product with little or no preparation or cleaning expense. The need for some economical method for recovering coal from thin seams is becoming more pressing as the more profitably mined seams are exhausted. Basic Principles of Auger Mining The principles of the auger, or screw conveyor, have long been used in drilling small diameter holes for blasting and other purposes. These same fundamentals apply to large diameter drilling. However, considerably more attention was given to designing a cutting head that would not pulverize the coal and consequently would require a minimum of power to drive it. The head, as designed for this job, consists of a cylinder 25 in. in diam with steel bits set in 3 positions spaced around the forward rim. A conventional coal drill bit, at the head of a bursting cone in the center of the cylinder, trails the forward rim slightly, see Fig. 1. The bursting cone and auger connections are anchored to the cutter barrel by 3 arms spaced at 120". This head design proved to be very satis-

Jan 1, 1953

By J. P. Newell, R. W. Storey

This paper describes the development of a continuous float-and-sink process to produce coal low enough in ash content to be suitable for production of electrodes. The cleaned coal had a combined iron and silicon content of 0.239 pct. The efficiency of recovery was 84 pct. AUGER mining is a form of continuous mining in that it completely replaces with a one-cycle operation the older conventional cut, drill, shoot, and load method of mining. Relatively new, having been utilized only within the last few years, it is even more continuous in its production of coal than some other forms of continuous mining being practiced in this country today. There are at present coal recovery augers in operation working along strip-mined highwalls from the outside. This paper will deal with underground mining, specifically with an experiment which was carried out with a machine developed for this purpose. The location of this experiment was in Pike County in eastern Kentucky in the upper split of the No. 3 Elkhorn seam. Coal height was 32 in. and roof conditions were very bad, with 10 in. of soft draw slate coming down with every cut in conventional mining. The coal is high-grade bituminous, excellent for metallurgical coking. Underground auger mining fills a need for a method of recovering coal in thin seams, under bad roof conditions or in split seams. In many cases it might save money by eliminating the handling of excessive quantities of rock. In fact, if the coal can be recovered, as in the case of this experiment, without mixing with any impurities outside the seam, it may be possible to load or ship a satisfactory product with little or no preparation or cleaning expense. The need for some economical method for recovering coal from thin seams is becoming more pressing as the more profitably mined seams are exhausted. Basic Principles of Auger Mining The principles of the auger, or screw conveyor, have long been used in drilling small diameter holes for blasting and other purposes. These same fundamentals apply to large diameter drilling. However, considerably more attention was given to designing a cutting head that would not pulverize the coal and consequently would require a minimum of power to drive it. The head, as designed for this job, consists of a cylinder 25 in. in diam with steel bits set in 3 positions spaced around the forward rim. A conventional coal drill bit, at the head of a bursting cone in the center of the cylinder, trails the forward rim slightly, see Fig. 1. The bursting cone and auger connections are anchored to the cutter barrel by 3 arms spaced at 120". This head design proved to be very satis-

Jan 1, 1953

By Luis Toron, Salvador Cortes-Obregon

John D. Price (Colorado Fuel and Iron Corp., Pueblo, Colo)—The paper on the coal fields of the Oaxaca district as prepared by engineers Toron and Cortes-Obregon of the staff of the Bank of Mexico bears witness to the thorough and careful way in which the men associated with this organization perform their work. There is little to be added to their paper in way of discussion other than to confirm and amplify some of their statements. Since the only extensive and well-developed field of coking coal lies in the northeastern section of the country adjacent to Sabinas in the state of Coahuila, it follows that blast furnace plants would be located in that same region. Two such plants are now operating at Monterrey and Monclova, using coke produced at the Sabinas district mines. But the nearer of these two plants is 600 miles from Mexico City and even farther from the center of population. Transportation of products from these mills to the market area is therefore expensive, both because of the distance and the difficulty of the terrain over which it must be carried. The development of an integrated steel industry closer to the center of population has therefore long been a goal toward which the Mexican technicians have been striving. While the presence of coal of some grade has been reported in many of the states, and many ideas have been advanced regarding its possible uses in iron and steel production, deposits of anthracite in Sonora and the various coals of the Oaxaca district as reported on in this paper are the only ones that have been explored in a serious manner. The coking coal from the Mix-tepec zone appears to offer promise of producing a coke which could be used in a standard blast furnace. Several problems are indicated, however: 1—The ash in the coal is high as mined, but indications are that it can be washed to an ash content of 15 pct with a recovery of 70 pct of washed coal. 2—Such washing would increase the volatile content from 17.4 pct to about 20 pct, and in a byproduct oven this should give a coke yield of close to 80 pct with an ash content of coke under 20 pct. 3—A free swelling index of 5 appears low for a good coking coal, and below that of the coals from the Sabinas district, which show between 6 and 9. But washing of the coal should result in an improvement in this regard; in the United States coals from Utah with an index even lower than 5 have made a usable coke. 4—A coal with volatile as low as 17.4 in raw coal and 20 in washed coal would come close to being classed as a low-volatile rather than medium-volatile coal, and low-volatile coals are notorious for their high expansion properties. Several plants in the United States are making coke from straight medium-volatile coal of 26 to 28 volatile content, and one at Rosita,, Mexico, from coal of 25 volatile. But no plants to my knowledge are using coal as low as 20 volatile. Since the Rosita coal appears to be a borderline coal from the angle of its expansion properties the coking of one of the straight lower volatile must be approached with caution. 5—There are few coals possessing any degree of coking properties which cannot be used in coke production by careful attention to its preparation and blending. The fact that coals of other types are available in this same region make improvement through blending very possible. 6—There are other workable methods of reducing iron ore other than the conventional coke-blast-furnace method. These will not be discussed here but it is known that their use has been considered. The technicians of not only Mexico but also of the other Latin American countries are keenly aware of their natural resources and their national needs. This paper emphasizes the fact that the Mexican technicians are working on their problem and attempting to speed the day of self-sufficiency for their country. Salvatore Cortes-Obregon (author's reply)—I wish to thank Mr. Price for his kind remarks. The Mixtepec coal as shown in Table II has 30 pct ash and a free swelling index of 5, but when the same coal is washed to 15 pct ash it has a free swelling index of 8 to 9 and the volatiles increased from 17.4 to 20.7 pct. A satisfactory coke has been produced from blends made in the Mexican laboratory using at least 40 pct of the Mixtepec coking coals with the other Oaxaca non-coking coals. Koppers in Germany report good coke obtained from the Oaxaca coal with a blend of 80 pct Mixtepec coal. Consideration is being given the possibility of using methods other than the conventional blast furnace for the reduction of iron ore near the Oaxaca area; electric furnaces appear promising. The non-coking coals could be used to produce cheap electric energy and the coking coals to make metallurgical coke.

Jan 1, 1955

By J. D. Kalasky

Continuous mining has revolutionized the coal industry but intensified the problems of earlier mechanization. From the installation of the first miner, it was recognized that face ventilation would be one of these problems. To establish an efficient work cycle for the equipment being used, conventional mechanized mining provides four or more development faces. Continuous mining, relying on a single machine, requires only one face, although more than one can be driven alternately by the same machine. In gaseous areas mined by the older methods, the intermittent rate of exposure allows time for the face to bleed off or be cleared, and the width obtained with conventional units makes it possible to hang a brattice line almost to the face. Continuous mining increases the ventilation hazards; there is no margin of time, and very limited space. For all practical purposes, continuous miners available today can be grouped into two major types, rippers and borers. The two can be further classified, respectively, as flexible and rigid. Both types have been in operation in U. S. coal mines for the past five to ten years in a number of seams, under various conditions ranging from slightly gaseous to very gaseous. Common methods of face ventilation utilizing jute or plastic brattice material for line curtains, stoppings, and checks have so far met with success. Air volumes to operating sections are greater than for conventional mining, and specific instructions have been given to section crews stressing the precautions needed in dealing with a steady rate of gas emission and the importance of maintaining reasonably air-tight brattice to direct a good steady flow toward the face at all times. Operators of continuous miners, as well as the foreman on each section, are equipped with an approved safety lamp for gas testing. When the miner is in steady operation, tests for gas at the face are made at 15 to 30-min intervals, depending on the

Jan 1, 1960

By D. W. Mitchell

J. H. Schlobohm (Joy Manufacturing Co., New York) —This paper has been read with a good deal of interest; however, there are several salient features which Mr. Mitchell has overlooked. The initial emphasis on selection should be placed on type of material to be loaded and method of mining involved. There is no possible connection in the comparison of scraper or Joy Lohite loaders with loaders of any other type. Neither can an operating comparison be made between "Rocker type" loaders and those equipped with loading head and gathering arms. Each unit has its application, but the application depends on material and the mining system involved. Mr. Mitchell's compilation of data on various loaders, while it represents a great many facts, does not represent true operational data, i.e. Table I, column heading "Max. Rated Capacity for 7 hr;" these figures are meaningless inasmuch as little or no time at all is allowed for tramming or switching or moving the loader from place to place. Table I, column heading, "Best Reported Field Shift," would be a better source of operational data, assuming of course that conditions, applicable to each class of loaders, were average to avoid enthusiastic performance figures. Table I, column heading, "Repair Parts Cost per Ton," is still another item which will not stand up under comparison unless the loaders involved are put into type classes, i.e., there are simply no means possible to show the repair costs of scraper loading versus a caterpillar mounted loader in abrasive material. The source of information here is also important. Actual field tests have proved that these costs are of a constant nature. Manufacturers are striving to hold them to a minimum and many interesting studies have been made. A fairly recent example of this in a large United States gold producer has proved that Joy HL-3 loaders can be consistently average, less than $0.025 per ton mucked. This cost represents repair labor and parts over a two year

Jan 1, 1952

By A. M. Keenan, R. H. Carpenter

Geologic faults have always been a plague to the mining industry, and have often reduced a mining venture from a profitable to a marginal operation, and even at times have forced companies to liquidation or bankruptcy. The use of geology as an economic tool in coal mining is not generally recognized, particularly in coal deposits which are horizontal or essentially flat-lying, in which faulting might not be prevalent. However in pitching coal seams, the existence of faults is considered an inherent challenge, as the inclination of the seam or seams per se, is a product of geologic disturbance. The problems of faults are a constant challenge to the management of the Thompson Creek Coal and Coke Corp. in their operations in Pitkin County, Colorado, where high grade medium volatile coking coal is mined. The background for this paper has been provided by the experiences of these operations, in overcoming the effects of faults in mining, and resolving the fault structures. The properties of the Thompson Creek Coal and Coke Corp. are located in central Colorado, 13 miles southwest of Carbondale, Colorado, a tow] on the main highway between Glenwood Springs and Aspen, Fig. 1. The elevation at the mine portals is approximately 8000 feet: there is a relief of some 2.000 feet to the adjacent ridges rising away from the [nine portals. 'The entire present production of these properties totaling 180.000 tons of clean coal per year is trucked to the railhead at Carbondale, Colorado, from whence it is shipped to the Geneva Steel Plant at Provo, Utah. The properties were initially opened in September 1952. A total of 125 men are currently employed with the production conting from the Number 1 and 3 Mines, and development workings of the Number 2 and 5 Mines respectively. While the No. 2 Vine was opened in 1953 prior to No, 3 Mine, the encountering of faults without sufficient development work accomplished to sustain production, necessitated the diversion of manpower to the other [nines for economic considerations. GEOLOGY '1-he productive coal seams occur in the Mesaverde formation of Late Cretaceous age which outcrop along the eastern edge of the Piance Creek Basin from the Grand Hogback, a part of the southwestern flank of the White River Uplift north of Glenwood Springs, southward through the 'Thompson Creek area and the

Jan 1, 1961

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 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

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