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By Paul A. Taylor

MUCH information concerning pegmatites which was thought to be true a few years ago has been proved false, and what is now actually known about some pegmatites is not true of many others. The erratic and seemingly unpredictable structure and variable composition of this class of mineral deposits has been widely emphasized. Even parts of the same pegmatite body exhibit marked differences in texture, mineral composition, width, and attitude. Constructive geological thinking in respect to pegmatites now aims to establish general laws rather than to stress the confusing diversity of features having no special economic significance. Substantial progress has been made in classifying different types of these deposits according to general features, internal structure, mineralogy, and origin. In some cases it has even been possible to block out tonnage reserves in advance of mining. It is still easy, however, to make highly erroneous predictions after a preliminary examination of a pegmatite prospect. Pegmatites are important to the economic well being of the country and to its military security. They furnish virtually all the feldspar, strategic mica, beryl, columbium, tantalum, and caesium utilized in the United States, as well as sundry other minerals and significant amounts of lithium and rare earth minerals and gems. With the exception of vermiculite, occasional ilmenite-rutile, and perhaps soda-lime feldspar and garnet deposits, basic pegmatites are of little economic importance. Consequently in this paper, as in common parlance, the term pegmatite generally relates to coarsegrained acidic rocks or what is aptly called giant granite. Available data indicating the size and importance of the production and trade in specified pegmatite minerals are summarized in Table I. Geological Features Much of the latest thinking on the economic geology of pegmatites is now available in a 115-page monograph' by a group of experts who participated with geologists of the Federal Geological Survey in the widespread wartime investigations. Doubtless the most significant feature of the monograph is indicated by the title, The Internal Structure of Pegmatites, but it also contains a vast amount of other new information and includes the assimilated concepts of many earlier writers, whose works are given in a comprehensive list of references. The shape, size, attitude, and continuity of many pegmatite bodies is controlled by the structure of the older rocks in which they occur. If the older rocks are easily penetrated, e.g., biotite schist, most of the pegmatites in a given district will be found outside the parent granite mass as exterior pegmatites. Marginal pegmatites are more prevalent if the older rocks are massive, unsheared, and sparingly jointed. Networks of pegmatites are abundant in highly-jointed rocks. In strongly foliated schists the bodies are usually lenticular, whereas in highly-folded areas they assume tear drop, pipe or pod-like, bow-shaped, or sinuous forms. Jahns2 recognizes five major shape classes: l—dikes, sills, pipes, and elongate pods; 2— dikes, sills, pipes, and pods with bends, protuberances, or other irregularities; 3-—trough-or scoop-shaped bodies with or without complicating branches; 4—bodies with the form of an inverted trough or scoop; and 5—other bodies, including combinations of the above and miscellaneous shapes. Many pegmatite deposits are scarcely big enough to be recognizable as such. Most of them, in fact, are small tabular deposits less than 4 in. wide and usually without economic concentrations of minerals. On the other hand, some pegmatites are more than a mile long and over 500 ft wide. The ratios of length to breadth range from 1 : 1 to 1 : 100. Although the vertical dimension bears no invariable relationship to strike length, tabular deposits or large lenses are often symmetrical enough to show nearly as much continuity down dip as in their horizontal extension, and some pipes or pods are amazingly persistent in the vertical plane. Small pegmatites often string along a fairly definite trend line; in a given district major bodies may lie roughly parallel, and where only a few of them do not, the erratically disposed bodies generally differ in composition from those conforming to the regular pattern. This does not apply, however, in all districts. Characteristically, pegmatite veins pinch and swell or split into branches. When they pinch out entirely it is often possible to find a new body by prospecting the extension of the strike or dip, but the chances of finding a hidden deposit are ordinarily too uncertain to justify much subsurface prospecting. Diamond drilling may yield valuable information as to the continuity of known deposits whose upper portions are well-exposed. Some deposits, in fact, can be proved up for hundreds of feet by surface trenching and then intersected by drill holes at various depths like any other vein-like deposit. Others twist and branch, apparently defying all efforts to explore them short of actual mining.

Jan 1, 1954

By F. G. Miller

Earlier studies of the mechanism of flotation showed that coal flotation products are contaminated by composite coal and mineral particles. The mineral matter of these particles has the dual effect of increasing the specific gravity and decreasing the coal surface of the particle available for bubble attachment. The present study showed that these particles are more loosely held in the froth column than the high grade coal particles, especially in the coarser sizes. Consequently, it is possible in the laboratory to produce a secondary concentration of the product by sprinkling the froth product with water to wash these loosely held particles from the froth during formation and recovery. To improve coal flotation, Bethlehem Steel Corp.'s research department has been working on a number of related problems based on both physical and chemical factors. The ideal is of course full understanding of the interrelated effects of both sets of factors. However, detailed studies1 of each physical factor have produced immediately usable results in commercial coal flotation and are providing the basis for our current work on the effect of chemical factors. Since these detailed studies provided the basis for the present investigation of froth sprinkling, their key results are highlighted here. For example, for a highly floatable bituminous coal, particle floatability was shown to be directly related to the percentage of coal surface available for bubble attachment and inversely related to particle mass. The data also underscored that in a given size range the addition of mineral matter to a particle makes flotation more difficult both by adding to the mass of the particle and by decreasing the amount of coal surface available for bubble attachment. Size also affects particle floatability in that as the size of particles of a given specific gravity increases, mass increases and the ratio between coal surface area and mass of the particle decreases. Flotation tests showed that any recovery of the coarsest particles in a coal feed is accompanied by progressively higher recovery of finer and finer particles. Consequently, fine high-specific-gravity particles are recovered with the coarse good coal. As a result, when the entire 14 mesh x 0 size range is processed by froth flotation, the good coarse coal is contaminated with high-specific-gravity fines. It is the small quantities of these fines that must be rejected from the froth product if flotation efficiency is to be increased above its present level. These principles and findings, together with work by Professor V. I. Klassen,' led to the froth sprinkling investigation reported in this paper. Working with Russian coals, Klassen had shown that froth sprinkling washes loosely held refuse particles from the froth and compensates for the water lost in the upper froth layers and increases its stability. As a result, the coal particles are retained in the froth and the loss of coarse coal from the froth is prevented. Decreases in ash of 0.1% to 1.5% were reported by Klassen. Working with a 28 mesh x 0 American coal, the author of the present paper found that single-stage flotation coupled with froth sprinkling produced a product comparable to that obtained by the more expensive method of roughing and cleaning.3 The investigation reported in the present paper expanded study of the effect of froth sprinkling to include each of the standard size fractions of a bituminous coal. The main findings are as follows: (1) For those size fractions between 48 mesh and 150 mesh, froth sprinkling produces a marked reduction in both sulfur and ash. In the case of those sizes coarser than 48 mesh, while this reduction is not as significant, the additional benefit of shorter retention time is still a plus for froth sprinkling. Furthermore, when froth sprinkling is coupled with normal fast flotation, the product purity approaches that obtained by starvation reagent feeding, i.e., gradual feeding of flotation reagent to float the recoverable coal from the pulp slowly in several increments. The specific results for these two size groups are: (a) 14 x 48 Mesh. By using froth sprinkling with fast flotation, product sulfur can be reduced to a level

Jan 1, 1970

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 J. R. Lucas, H. E. Rutherford

The primary objective of this research was to evaluate the feasibility of ultrasonic testing for the purpose of analyzing the structure and properties of rock material. A secondary objective of this study was to determine experimentally the elastic wnstants of various rock specimens with the use of ultrasonic equipment. Elastic constants can definitely be determined quickly and accurately by ultrasonic investigation. With proper equipment there are indications that ultrasonic waves could be utilized in the detection of flaws, such as fractures and planes of weakness, measurement of thickness, and the study of structure of rock material. Ultrasonic principles have been applied in many industries for the purpose of determining the structural characteristics of various materials. The principal applications of interest here are detecting flaws, measuring thicknesses, and determining elastic constants. The main advantage of ultrasonic analysis is that it is nondestructive. It also has extreme sensitivity, high penetrability, good accuracy, rapid response, and the necessity to have access to only one side of the specimen. These various applications and advantages lead one to think that ultrasonic principles might be applied to rock material in situ for purposes of analyzing its structural characteristics. Therefore, the primary objective of this research was to evaluate the feasibility of ultrasonic testing for the purpose of analyzing the structure and properties of rock material. A secondary objective of this research was to determine experimentally the elastic constants of various rock specimens with the use of ultrasonic equipment. The term ultrasonics is given to that branch of acoustics dealing with periodic sounds whose fre- quencies are above the audible range or above approximately 15,.000 to 20,000 cps. The first practical production of this high frequency sound was made possible after World War I due to developments of electronically controlled sound wave sources. At the present, the maximum practically attainable ultrasonic frequency is approximately 1000 megacycles per sec. Sound waves of sufficiently high frequency can be made to travel in a beam with little spreading. Therefore, such a beam can explore a material and detect nonhomogeneities in it by reflection. This is the principle of the ultrasonic reflectoscope invented by F. A. Firestone' in 1942. An ultrasonic beam, on pulse-penetratink a solid material medium, produces an echo when it strikes a flaw which has different acoustical properties than the surrounding medium. This echo can be detected by the same transducer which sends out the original pulse. This principle is the basis of the important industrial techniques of nondestructive ultrasonic testing of metals and other solid material. Later developments in the field of ultrasonic testing of solid material have been mostly improvements in equipment and instrumentation. To date, most of the industrial applications of ultrasonic analysis of solid bodies have been in the aircraft, electrical, and nuclear energy fields on the nondestructive testing of metal. The one solid material similar somewhat to rock material, upon which some work has been done to apply ultrasonic testing techniques to study structural integrity, is concrete. Jones and Gatfield3 briefly summarize this work on concrete. Their basic conclusion was that the strength of concrete was a function of the change of the longitudinal sound wave velocity in the concrete. Only recently has any extensive study been made to apply ultrasonic testing techniques to rock material. Lutsch studied rock material by an ultrasonic pulse method to determine the extent and degree of fracture of rock faces. Lutsch states that the number of observed reflections is a measure of the degree of fracture and that only ultrasonic waves are able to detect fractured zones subjected to high pressure. Suzuki, Sasaki, and Siohara5 studied the characteristics of the sonic wave in rock for propagation over a wide range of wave lengths. They indicated that, in rock, the propagation constant and attenuation of the

Jan 1, 1962

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 A. C. Watts

For several years, a mine in Colorado experienced considerable trouble from small fires caused by the blasting of coal. Although a well-known make of permissible powder was used, it was first thought that the powder ignited gas that might be present and that this, in turn, set fire to line brattices or to the coal. Experiments were, therefore, conducted with another make of powder, but the results convinced the superintendent of the mine, Robt. Williams, Jr., that the fires were not caused by the powder but by the high-voltage electric current used for detonating the shots. The mine "makes" gas very freely, giving off from three-fourths to one million cubic feet of methane every 24 hours. The coal is a good grade bituminous of a very free burning quality. Electric shot-firing from the outside, after all men and animals were outside, had been practiced for a number of years. The shooting circuit was connected with the main power circuit at the mouth of each entry. Shooting circuit switches were mounted in boxes, which were kept locked with the handle of the switch, in open position, showing through the bottom of the box. The door of the box could not be shut and locked when the shooting switch was closed. At the entrance to each room, or working place, a small switch, called the miner's switch, was placed on the shooting line. It was the duty of the miner, when entering his place at the beginning of the shift, to see that this switch was open and that it was kept open until he left the placc at the end of the shift to go out of the mine. Grounded return was used for main power lines. The current was 500 volts, direct current, with 100 to 300 amp. The nearest working face was 9600 ft, from the shot-firers' switch outside the mine and the most remote working face was 2½ miles. The various working faces were scattered over a wide area between these two points. A description of the main power line would be valueless; the conductors were of sufficient size, the drop in voltage being not more than

Jan 1, 1927

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 W. N. Poundstone

The mining and ventilating system used in development work in the Pittsburgh Seam in northern West Virginia is discussed. The seam conditions and the nature of the accompanying methane gas are described. The type of equipment and the mining cycle will be discussed, showing how they are well suited for very gaseous development work. Face ventilation in development work is possibly the fastest growing problem of the industry. The coal mines of the future will be faced with the prospect of mining from under increasing depths of cover. Consequently, larger and larger amounts of methane gas probably will be found. The Pittsburgh Seam, in northern West Virginia, is an example of an area already faced with this problem. At the present time, most of the development work being done in this seam lies beneath 500 to 1200 ft of cover. The Pittsburgh Seam in this area has always been very gassy—even near the outcrop—and the recent development work has been accompanied with extremely large volumes of gas. In many cases, a single development section has liberated in excess of 1,000,000 cfm in 24 hr. This problem of heavy gas liberation was the chief concern, several years ago, when continuous mining equipment was first considered at Christopher Coal Co. All of us were apprehensive about the liberation that would accompany rapid extraction in a single working place. However, the experience during the past few years has shown that this ability to mine only one place at a time, is actually the key to working this type of coal. With all of the mining or advancement concentrated in one place, the ventilation can also be concentrated. By this it is meant that continuous mining permits the active working place to be ventilated with a maximum amount of fresh air, taken directly from the intake source without first passing another working place. Continuous mining (and a good ventilating system) also permits a much greater concentration of attention or vigilance to the actual working place. There are two things that are very important to the mining of coal having high rates of liberation. First, adequate volumes of air are necessary. Second, and perhaps more important, a mining and ventilating system must be used that will provide an uninterrupted flow of air to every portion of the working section. Liberations of this magnitude take only a few seconds of interruption for a dangerous accumulation of gas to occur. With adequate volumes of air available, the ability to concentrate ventilation more than offsets the concentration of gas emission that is inherent with continuous mining. The mining system used with continuous mining equipment at Humphrey Mine is similar to the system used at many mines in the area for development work. This system is designed to favor ventilation, realizing that other efficiencies are meaningless if the equipment must stop because of ventilation difficulties. This plan is especially well suited to minimizing ventilation interruptions. Basically, the overall plan of mining is to develop headings into virgin coal and encircle or block out large areas. These blocks are generally at least 2000 ft sq. The purpose of this blocking out is to bleed gas from the area before pillaring. Experience has shown this method to be quite effective, even in the most gassy areas. The gas in this field seems to migrate or flow readily from the solid coal into the outside return headings of the development work. The numerous clay veins and slips that are found in the area are extremely good avenues for gas flow. A block of coal, surrounded with development headings, usually bleeds off readily; and since it is cut off from the virgin coal, it is not subject to gas migration, through the seam, from this source. However, the outside return of the encircling development work, adjacent to the virgin coal, may liberate gas for years. This liberation from the outside ribs in the virgin coal is the reason split ventilation is used in development work. If split ventilation were not used, there would, in many cases, be a serious build-up of gas in the intake before it could reach the working face. Fig. 1 shows a typical development section having seven headings. The two outside places on each side are returns, and the three center headings serve as intakes. This section is equipped with a ripper-type continuous mining machine. An off-track loading machine is used to load from a surge pile on the

Jan 1, 1961

By D. W. Mitchell

INE operators have a choice of several classi- fications of mechanical loaders. Within each classification there are many types and makes available. Table I lists loaders on which manufacturing data as to operating characteristics are available. This paper discusses the conditions met in a mine as they affect these characteristics. It is assumed that management will provide satisfactory engineering, supervision, power, maintenance, loader service within a concentration of workings, and a balanced working cycle served by balanced production equipment. These are factors which affect the optimum operational efficiency of a loader. The conditions which determine loader choice are: Height of Vein—The main limitation to the use of a loading machine is the height to which the vein is mined; i.e. a loader cannot be higher than the place in which it is to work unless rock is taken. Since there are several low vein loaders being developed and successfully applied, this does not appear to be a requisite of an efficient mining operation. The maximum useable height of a loader should be equal to the working seam thickness less a working clearance for travel and operation. A—Working seam thickness = B—Artificial roof support thickness = C—Safe headroom below support = (generally about 6 in.) D—Height of roadway = E—Total of B + C f D =_______________________ A—E = maximum loader height F—Distance from top of roadway to bottom of roof support = G—-Height of loader's coal or ore line (see Table I) = H—Difference between F and G = Size of largest broken particle + a safety factor of 2 to 3 in. H is a loadability factor and is important because the clearance over the top of the conveyor chain to the roof or roof support must be greater than the largest broken particle or else the particle and/or the conveyor chain may be broken. From this standpoint it might be desirable to use a lower height machine even though it may have a smaller capacity, though this is warranted only when increased lump realization offsets lower man-day and machine pro- duction rates. Ivan Given has stated, "Loading should not be handicapped by securing lump only to grind it in the breaker." Width of Working Place—The width of a working place is the greatest width a place may be safely driven. The working place includes not only the face but the roads the loader must tram when mucking out more than one face. The width is limited by the system of placing props with sufficient room between them for the machine runner (about 2 ft). Maximum width of place in which the loader will operate at maximum efficiency is determined by the type of mounting. With track-mounted loaders, maximum width of place is determined by maximum angular swing of the loading head and proximity of track to the face. The use of double track leads to added expense and operational complexities. Several mines that have used double-track systems have found that a substantial saving is made by the use of machines that load out wider places. Trackless loaders may work in rooms of any maximum width. Minimum width of place in which a loader will operate is determined by width of the loader plus a safe movement area for the operator. Often a barodynamic study of the mine will show that either an increase in room width and/or a change in the system of propping may be made by using roof bolting, full or part-width room timbers, removable aluminum I-beams at the face, etc. Increasing the width of a working place will, in general, increase loader operating efficiency by providing more material per fall and by decreasing the preparation and moving time per ton. A change in the system of propping that would increase the minimum width might permit the use of a higher capacity loader with safer working conditions. Proper face tirnbering is necessary to insure safety from roof falls and because the speed of preparation and loading generally increases since the men work with greater assurance and there is less chance of kicking out posts. Maximum Room Length—With the exception of the scraper and the duckbill, the limit to room

Jan 1, 1952

By Orville R. Lyons

THE continuous vacuum filter, whether it be a vertical or horizontal disk, inside or outside drum, when properly applied is a very effective device for dewatering fine solids. Because of the number of variables that affect its performance, however, it is not possible to obtain consistently maximum moisture reduction in an operating plant. This paper presents a method for determining the average cake moisture and the variations in cake moisture to be anticipated when dewatering fine solids by continuous vacuum filters under normal operating conditions. Previous investigators1,2 have found that certain factors influence filter performance, drum or disk diameter, speed of rotation, percentage of immersion, and the percentage of solids in the feed influence capacity. Blinding of the filter cloth reduces effective vacuum, produces a wetter cake, and may decrease capacity. Reduced vacuum or improper adjustment between vacuum and blowing portions of the cycle will produce a wetter cake and also may decrease capacity. Widely fluctuating feed tonnages will adversely affect cake moisture and capacity. Flocculation, when properly applied, increases capacity and reduces cake moisture. Without flocculation, the slimes are the first particles sucked to the filter medium, which becomes partially blinded in consequence. With flocculation, the flocs act as coarse particles and the filter medium is not blinded. Overflocculated pulps, although filtering well, may yield wet cakes. Average Particle Size and Cake Moisture A method for predetermining the surface moisture content of coal filter cakes was published in 1949; but this method could not be applied to coals deficient in —200 mesh material or to fine mineral concentrates. Test data obtained for geared-weight types of vibrating screens4 and for centrifuges5 used to de-water coal provided definite correlations between the average particle size of a dewatered cake and the surface moisture content of the cake. It was thought that this same method of correlation should be applicable to filters, whether dewatering fine coal or fine mineral concentrates. To develop a correlation or correlations between average size of cake and cake moisture, the author collected data for filtration operations from as many sources as possible. In a considerable number of instances the information was provided by plant preparation engineers,"-" a few of the tests were on a laboratory scale: and the remainder of the data was obtained from published articles.10-17 Table I contains size-consist and miscellaneous data for coal filter cakes, and Table II contains the same data for mineral concentrate filter cakes. Fig. 1 shows the relationships existing between the average particle size of the cake and cake moisture (surface moisture for coal and total moisture for minerals) when dewatering fine solids by a variety of makes and types of filters. Average particle size is calculated by averaging the linear openings of the various pairs of screens used in the screening test, weighted in accordance with the percentages of material in the various fractions.

Jan 1, 1952

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 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 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