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By H. Rush Spedden, Arvid Thunaes

Pilot plant test work in 1942 and 1943 showed that by a combination of deslim-ing, fine-size classification, and Sullivan deck concentration it is possible to recover heavy minerals such as cassiterite at least as fine as 10 microns in size. This appreciable improvement in gravity concentration practice has been substantiated by several full-sized plants. IN the past, mills treating ores of tin and tungsten by gravity concentration have recovered very little mineral finer than 325 mesh, although some form of slime concentration has been generally attempted by the use of buddies or round tables. This paper describes a series of pilot plant tests made in 1942 and 1943 in which the use of the Sullivan deck was investigated for the recovery of an appreciable amount of the fine values formerly lost. The investigation was initiated by the U. S. Government in an effort to increase the wartime production of tin. Bolivian milling practice at that time included jigging of the coarse sizes, tabling of the intermediate sizes on conventional shaking tables, and the use of buddies or round tables on the finest sizes. Flotation of the pyrite was employed at different stages of the treatment, depending upon the quantity present and the preference of the operator. Usually good recoveries and the bulk of the production were made in the jigging and tabling operations provided that these sections were not overloaded. Buddies and round tables were used for a small additional recovery if it could be shown that they could pay for their high cost of operation. Extensive test work established that by a combination of desliming, classification, and Sullivan deck treatment, cassiterite as fine as 10 microns (1500 mesh) could be recovered when proper conditions were maintained. This test work has been verified by several full-scale installations. The method is generally applicable to the preconcentra-tion of large tonnages of low-grade ores or tailings for the recovery of a small amount of a valuable heavy mineral. The Sullivan deck (or now the Denver Buckman tilting concentrator), which is the essential part of the method, was devised originally for the recovery of tin from tailings of the Sullivan concentrator, Kimberly, B. C. Since the material treated contained only 0.05 pct Sn, it was impractical to attempt to recover cassiterite finer than 500 mesh.*

Jan 1, 1951

By Benjamin Gildersleeve

Uniquely colored, thin-bedded quartzite is quarried between Crossville and Crab Orchard in Cumberland County, Tenn. It is produced in all sizes up to the limits of transportation from beds usually ranging in thickness from less than Vi in. to 6 in. The stone is very resistant to wear and weathering and is marketed throughout the country for many different purposes in exterior and interior construction. The annual production exceeds 36,000 tons, valued at nearly $1,000,000. THE building stone of the Crab Orchard District is a fine-grained, thin-bedded sandstone which is so strongly cemented by silica that it may be classed as quartzite. As used in this paper the term building stone includes dimension stone, rubble, ashlar, flagstone, and rough broken stone. The Crab Orchard District is in the central portion of Cumberland County, Tenn. It is named from the town of Crab Orchard near which the stone was quarried first for industrial use. The heart of the district is an area of some 1200 acres from 1 to 3 miles east of Crossville which is 70 miles west of . Knoxville and 125 miles east of Nashville (fig. 1). At the present time, there are about 15 quarries worked by companies and individuals. Geology: The Crab Orchard District is on the Cumberland Plateau, or Cumberland Mountain as it is called frequently. Most of the area is characterized by a comparatively even surface from 1800 to 2000 ft in elevation and is drained by streams flowing into the Tennessee River. Included in the district are two quarries located 7 miles west of Cross-ville and this area is drained by tributaries of the Cumberland River. Most of the quarries are located near the 1800-ft contour which is just below the general upland surface. The surface rock over most of the district is the Rockcastle sandstone which has a thickness of about 260 ft.1 It is the topmost of the six formations of

Jan 1, 1951

By Phillip E. La Moreaux

Fluoride, generally less than 0.5 ppm, is present in ground water from rocks of Paleozoic age and older, in northern and eastern Alabama. Some of the water-bearing formations in the Coastal Plain area of the State yield water with as much as 6.8 ppm fluoride. IN June 1940, the U. S. Geological Survey, in cooperation with the Geological Survey of Alabama, began a study of the ground-water resources of the part of Alabama where water is obtained from Cretaceous rocks. The purpose of the study was to determine the quality, quantity, occurrence, and availability of ground water in that area. These studies have been expanded to include ground-water investigations in the area of Tertiary rocks, or southern quarter of the State, and certain areas in northern Alabama. The first report issued on these studies was by C. W. Carlston,1 of the Alabama Geological Survey. In the present paper all references to ground-water data for the Cretaceous area are taken from this earlier publication. In 1945, the Dentistry Division of the Alabama Department of Public Health became interested in the correlation of tooth decay and mottled enamel with the chemical quality of ground water used for public supplies in the State. Through a cooperative arrangement between the State Department of Public Health and the Ground Water Division (now Branch) of the U. S. Geological Survey, a report by the author2 was published in 1948. These two reports give accurate information on the occurrence of fluoride in the Coastal Plain of Alabama. At present, only scattered information is available on fluoride in ground water of the crystalline-rock or Piedmont area and for the Paleozoic area of Alabama, but it is hoped that in the future more complete information on the presence of fluoride in ground water in these areas can be collected. Outline of Geology and Ground Water: As described by Adams,3 the State of Alabama includes parts of two major geologic divisions, the Appalachian region and the Coastal Plain. The boundary between these divisions is irregular and is known as the Fall Line of the Atlantic and Gulf Coast States. The Fall Line enters Alabama near Phoenix City, extends westward to Wetumpka, Clanton, and Tuscaloosa, and then swings northwestward to the northwest corner of the State. The Appalachian region in Alabama includes three major provinces, the Piedmont province, the Appalachian Ridge and Valley province, and the Appalachian Plateau province (fig. 1). The rock formations in the Piedmont province in east-central Alabama are mainly of pre-Cambrian age, chiefly crystalline schists and gneisses injected by younger igneous rocks (fig. 1, I). They are faulted and folded and have a complicated structure. These rocks are the oldest and among the most complex rocks in the State. Generally, only small yields of ground water are obtained from rocks in this area. Even though ground water is of great importance for the development of domestic and farm supplies in rural areas and a few small industrial and municipal wells, yields in the area from individual wells generally range from 5 to 25 gpm and rarely exceed 50 gpm. The geologic formations of the Appalachian Ridge and Valley province and the Appalachian Plateau province (see fig. 1, II, and III) are separated on the basis of their structure. These rocks are of Paleozoic age, ranging from Cambrian to Carboniferous, and comprise a succession of formations consisting chiefly of shale, sandstone, limestone, and dolomite, aggregating many thousands of feet in

Jan 1, 1951

By R. Q. Shotts

Data are presented which show that fractions of varying densities from the same coals are oxidized at different rates by nitric acid. From oxidation data, the approximate quantity of "bright" and "dull" components may be calculated. Definite relationships between oxidation rate and rank are shown for density fractions from at least two bituminous coals. IN 1945, the Alabama State Mine Experiment Station initiated a program on the determination of fusain in certain coals of the state.1 Because of the lack of equipment and personnel necessary for microscopic examination, the chemical method2 was used exclusively. During studies of the distribution of fusain in fractions of various sizes and densities on a medium volatile coal, the author noted a higher rate of oxidation of the nonfusain part of the coal in the lower density fractions and the difference in rate became more marked the shorter the oxidation period used. Since durain and other "dull" components of coal have been reported3,4,5 to have higher densities than vitrain and the bright components, it appeared probable that the observed differences in rates of oxidation in the different density fractions were due to variations in petro-graphic composition. Such effects have been noticed by other workers.2,3 In order to obtain further evidence on this point, several Alabama coals of different rank have been separated into three or more fractions on the basis of density and their rates of oxidation determined, using the procedure specified for the determination of fusain, except that measurements were made at short time intervals as well as for the longer periods specified in the regular procedure. Significant differences in the rates of oxidation of the nonfusain part of the various fractions have been observed, and it is believed that studies of this type may make possible the identification of individual petrographic components or lead to methods for their semiquan-titative estimation. Properties of the Coals: Analyses and some properties of the whole coals studied are given in table I. The range in rank of the coals is not the widest possible within the state, but samples of the low volatile coals that occur in thin beds in the northeastern part of the state, were unavailable. The Clements bed is not, at present, considered workable. The coal is of medium volatile rank, quite friable, and very strongly swelling. Analyses of different size fractions show that the larger sizes are of higher rank than the smaller ones. Material passing the No. 50 sieve (all sieve sizes refer to the U. S. Standard Series) is close to high volatile A in rank, except for the size passing a No. 200 sieve, in which fusain is partially concentrated. Because of the variation in rank of the sizes of Clements coal, oxidation studies were made of two widely different size ranges. The + 1-in. material was crushed on rolls to about 1/8 in. and separated into five fractions by float-and-sink treatment in a benzene-carbon tetrachloride mixture. The — No. 4 + No. 8 and the — No. 4 + No. 50 size ranges tested overlap but the latter size was studied last with the expectation that it would include more oxidizable material than the first size. This proved not to be the case. Yields and properties of the specific gravity fractions oxidized are shown in table II. Milldale bed coal probably represents, as nearly as it is possible to do so, the "average" coal of the state. The material passing a No. 4 sieve was the only size range oxidized. It was divided into four specific gravity fractions. As in every case the heaviest fraction, which was essentially rock, was discarded. The Black Creek is the most extensive in area of

Jan 1, 1951

By B. T. Wykoff

This paper describes an important function of mining in Southeast Missouri. The practice is a necessary procedure to prevent disintegration of the pillars. Although the technique is peculiar to this district, it might have application in other mines using room and pillar method. CONSIDERING the subject alone, this paper would cover only a mechanical treatise of pillar roping. Because the St. Joseph Lead Co. originated the practice, it is proper to give a short introduction to the company. It is also proper to describe pillars and discuss their purpose to justify expending large sums of money to preserve them after the surrounding area has been thoroughly mined. Because the St. Joseph Lead Co. originated the practice of back pinning with wedge bolts, patch plates and channel irons, and the pouring of substitute pillars of concrete, these too, are mentioned. The Southeast Missouri Division of the St. Joseph Lead Co. is the largest lead producer in the United States. About 95 pct of the present producing mines are in St. Francois County, which begins 60 miles south of St. Louis. The entire lead ore production of this county, at the present time, is mined and milled by this company, treatment plant capacity being about 22,000 tons per day. This area is also the oldest, lead having been mined at Mine La Motte, in Madison County, in 1723. Estimated production of all the various companies that have mined here is about 7,000,000 short tons of pig lead. Method of Mining The lead occurs in the ore as galena. Galena is found either in the form of disseminated or solid streaks, of varying thickness. Ore stratum is the Bonne Terre dolomite, which is about 350 ft thick, but almost all the ore is in the lower 100 ft. Ore is mined by room and pillar method in open stopes. These stopes vary in height from 7 ft to 200 ft, depending on the thickness of the pay ore in that spot. Back, or roof, is supported entirely by pillars. In a majority of stopes, after the loose scales are mined down, the back has remained solid for many years (fig. 1). Under ideal conditions pillars are spaced to form an equilateral triangle and are as small and as far apart as the height of the stope and the character of the back will permit. They vary in diameter from 10 ft to 50 ft and are from 18 ft to 50 ft apart, measured from the outer edges. In general, the pillars are approximately 12 ft in diameter, and 25 ft apart. The spacing is very uncertain because of the vagaries of the ore trend, the desire to better support breaks in the back, and the attempt to spot out the thinner ore by not placing one in the core of the ore body. Use of Structural Steel Channels In stopes where the back is bad because the overlying formations are thin with wet, weak and shaly

Jan 1, 1951

By T. C. King

AT the new Graham-Central Mill of Eagle-Picher, near Galena, Ill., material is simultaneously dewatered and introduced into the ball-mill scoop boxes by the use of variable-speed, 24-in. spiral classifiers. This arrangement has proved satisfactory in that the moisture content of the wet feed is reduced, thus allowing maintenance of density within the mills, and a constant feed at any desired rate to the mills is provided. In this mill the primary feed is wet crushed to minus 1/2 in. and preconcentrated by jigs to produce a bulk lead-zinc-iron concentrate and a waste tailing. The jig concentrate is then dewatered by a conventional dewatering classifier and stored in bins ahead of the ball mills. Even though this material is well dewatered, the discharge at the bin bottom is sloppy because of the gradual drainage of the water from the surface of the particles. Conventional belt feeders with vertical cutoff gates are unsatisfactory because of the flushing characteristic of the wet sand. The new installation employs 24 in. x 10 ft, 9-in. Simplex double-pitch spiral classifiers for feeders from the bin bottoms to the ball mill scoop boxes (fig. 1). These units are arranged with straight tanks and are driven by 5 hp motors through a U. S. Vari-Drive to provide a 4 to 1 speed ratio. Each bin discharges through a chute into the rising side of the spiral. The flushing characteristic of the material from the bottom of the bin is utilized to maintain a steady feed. A small amount of water tapped into the bottom of the bin insures uniform feed to the classifier. Excess clear water drains over the weirs. Feed rate to the mills is controlled by the variable-speed drives. The feed delivered by the spiral at a given speed setting is relatively dry, uniform, and steady, and as such is easily controlled by the operator. The classifiers are set on a pitch of 5 in. per ft and have a sand-raking capacity of from 8 to 32 tons per hr. It is believed that the solution of this problem offers possibilities for the storage of deslimed rod-mill sands and other similar products of the industry.

Jan 1, 1951

By C. G. Fink, G. L. Putnam

The dissolution of gold by cyanide solutions was studied by determining the time required for the solvents to dissolve gold leaf. Minute traces, even 0.5 ppm, of sulphide ion retard the dissolution of gold, and this behavior cannot be accounted for by the presently accepted hypotheses involving oxygen-depletion or thiocyanate formation. On the other hand, traces of the salts of lead, bismuth, mercury, and thallium considerably accelerate the dissolution. The beneficial effect of lead is dependent upon the pH and cyanide ion concentration of the solution. ALTHOUGH cyanide solvents are used very ex-x*. tensively for the recovery of gold from ores and concentrates, the normal rate of solution of the precious metal as determined by Barsky, Swainson, and Hedley,1 and others2 is less than about 3 mg per sq cm of exposed gold surface per hour, corresponding to a corrosion depth of approximately 1.5 microns (0.00006 in.) per hr. Although the slow rate may be increased to some extent in certain cases, as, for example, when the the gold is partially imbedded in iron pyrites,3 the leaching operation is generally the most time-consuming step of the cyanide process. Frequently it is necessary for the leaching solvents to remain in contact with the ores for 50 to 200 hr in order to dissolve all of the gold particles.' Not only does the slow solution rate cause some increase in the capital costs, but the cyanide consumption also may be increased because of vaporization of hydrocyanic acid and secondary reactions with minerals such as malachite (CuCO3 • Cu(OH)2), and pyrrhotite, which often are associated with gold. It is not surprising that numerous attempts have been made to accelerate the dissolution of gold in cyanide solutions. Experimental Gold Leaf Test: The gold leaf test was apparently originated by M. Faraday, who, by means of it, discovered that dilute cyanide solutions would dissolve gold readily in the presence of oxygen. The usefulness and reliability of the test have been discussed elsewhere.5,6 Essentially, the method consists of determining the time required for 5-ml portions of cyanide solvents to dissolve squares of 23 carat gold leaf, 0.5 cm on edge, when shaken in 10-ml test tubes with the solvents. The weight of gold leaf per test was 0.051 mg as determined by the direct weighing of a leaf 8.6x8.6 cm (weight = 15 mg) and measuring the area of the test sample (0.25 sq cm). Gold leaf is of remarkably uniform thickness.6 Since gold leaf, like native gold, contains small amounts of copper and silver, the results are comparable with those one might expect in cyaniding gold ores. As six or more tests can be made simul- taneously, the relative efficiencies of the solvents can be evaluated readily. All dilute solutions of lead, bismuth, thallium and mercury salts were prepared by the ordinary dilution method used by chemists. Our C.P. sodium cyanide and buffering reagents had no detectable amounts of either sulphide ion or heavy metals. Soluble Sulphides in the Cyanide Process: The action of soluble sulphides is of interest in the study of the dissolution of gold in cyanide solutions. That gold ores which contain compounds of arsenic and antimony often are not amenable to direct cyanide treatment is well known, and it said that arsenic and antimony are "cyanicides."7 It is commonly believed that part of the sulphur content of such ores is soluble in alkaline solutions to give sulphide ion, which reacts with the free cyanide content of the leaching solutions to form inert thio-cyanates or reacts with the dissolved oxygen to form sulphites or sulphates.8 A purpose of this paper is to demonstrate that sulphide ion may behave in a manner that cannot be explained by either the oxygen-depletion or thiocyanate formation theories and to prove directly that concentrations of sulphide ion which cannot be detected even by colorimetric methods may retard the dissolution of gold.

Jan 1, 1951

By R. E. Zimmerman

Wide attention is being placed upon various methods for cleaning the fine sizes of bituminous coals. The author describes and analyzes the results achieved on wet concentrating tables of modern design on unclassified feeds of —Vz in. particle sizes, including tests on high-sulphur coals. THE increased economic necessity of cleaning the small sizes of coal has forced wide attention upon various methods or processes being used. Although numerous cleaning plants in the past have been cleaning coal smaller than Vz or % in., the large quantities of higher ash fines being produced by mechanical mining and the need for cleaner coal under competitive market conditions have brought about many new installations of fine-coal cleaning plants. Of various methods used, one of the more popular types is the wet concentrating table. Approximately 525 tables have been installed in the United States in the past four years of which 320 have been installed in 20, large new plants. There are now probably 1600 tables of all types in existence in this country for cleaning coal. The purpose of this paper is to describe the cleaning of fine coal by wet concentrating tables. Its scope is confined to a discussion of the theory of tabling and some of the factors entering into tabling action, a description of the Deister SuperDuty diagonal-deck coal table and a discussion of the results of tabling various fine sizes of coals. Al- though tables have been used for cleaning coal as coarse as 2 in. and favorable results are reported, this paper is limited to reporting on the results of cleaning coal fines from 1/2 in. down. Also, although it is well known that the performance of tables is even superior under classified or closely sized feed conditions, this paper concerns itself with the results achieved by tabling unclassified feeds. History of Wet Concentrating Tables The wet concentration of minerals by some form of tabling is quite old, starting originally from fixed, smooth surfaced, tilted boards taking advantage of the laminar flow of liquid films and the separation of minerals according to specific gravity. This was followed by the development of rough deck surfaces

Jan 1, 1951

By Arch J. Alexander

This paper is a resume of a study undertaken by the West Virginia Department of Mines. The underlying and direct causes of accidents are determined in each occupational group. Then from this study, accident prevention courses, visual and written, are to be taken to the various working groups in the mines. THE causes of accidents in West Virginia mines follow the main pattern of coal industry accidents in the United States—falls of roof, ribs and face, 56 pct; from haulage, 25 to 30 pet; blasting, 3 pet; electrocutions, 3 pct; explosions, 2 pct; and other causes, about 10 pct. To arrive at some satisfactory conclusions regarding causes and prevention of mine accidents, the Department of Mines has made a study of all mine accidents, from 1945 to 1948, inclusive. During the last two years of this period, we did not have a major explosion, one that took the lives of as many as five men. The direct cause of these accidents is apparent in the reports but we who are making a close study into their frequency think that the virus lies much deeper, and we want to suggest an underlying cause. The rapid growth of the coal industry in West Virginia, the largest producer of bituminous coal, has resulted in a demand for manpower which probably has never been equaled in mining history. At the turn of the century, West Virginia had 24,635 men employed in and about her mines, and they produced 21,153,341 tons of coal a year. In 1948, there were 125,669 men employed, and coal production for the year was 168,589,033 tons, an average increase of 35 pct for each year since 1900. Over 100,000,000 tons of this increase were mined in the southern counties of the state. A safety analyst might read into the history of this development an underlying cause for accident occurrence. He might say that in such a vast, rapid development the available supply of trained miners became exhausted and that young, inexperienced men were employed in the mines by the thousands. Being unacquainted with the hazards of mining, it would be only natural for these young men to be injured or killed in too great a number. This assumption would not be borne out by the records. Men were not injured or killed in disproportionate numbers during their first years in a mine. The only reason we can give is that, during these years, they possessed an attribute of safety—CAUTION—to a much greater degree than in the years following. The coal companies in West Virginia that have been operating over a number of years have a nucleus of employees who have been with them ten, twenty, thirty, and forty years; many of them having gained all of their mining experience with one company. Their habits of work have been

Jan 1, 1951

By Louis C. McCabe

The control of dusts and fumes of submicron size is involved in many process industries. This paper presents in tabular form the quantitative data from a number of metallurgical operations and discusses allowable discharge in terms of effluent-weight ratio. Typical installations for recovery of stack effluents are cited. THERE are many sources of smoke, fumes, gases, and dusts that may cause undesirable contamination in an industrial area. Fumes and gases are the result of chemical action or volatilization, while dusts are produced in processes of disintegration and dispersion. Commonly, smokes are suspended particles formed by combustion or destructive distillation. Methods for controlling smoke are generally known, and great improvements have been made in equipment for removing fine dusts and fumes from gases. In most of the industrial centers of the country, the chief duty of the air-pollution abatement agency is the elimination or reduction of emissions resulting from improper combustion. Correct methods of firing, proper fuel selection, and installation of properly designed equipment may accomplish this. It is becoming more widely recognized that much of the contamination of an industrial atmosphere is caused by dusts, fumes, and gases that are not related to the process of combustion. As such emissions are usually a functional part of plant operation, they rarely can be prevented even by better design and careful operation. When dusts, fumes, and gases are an integral part of the operation, mechanical equipment must be designed to remove these emissions. Most of the dusts and fumes that should be controlled are less than 1 micron in size. Control of such small particles is largely a new and untried field in many industrial areas. However, where conditions have been critical, effective control equipment for dusts, fumes, and gases has been in operation for many years. The problem is not one of smoke abatement and requires an entirely different set of standards for effective solution. Identical standards need not be applied to all industrial areas because of differences in meteorology, physiography, plant location, plant density, and nature of stack effluents. In this paper, discussion of control of atmospheric contamination by industrial plants refers specifically to the Los Angeles area, but the method of approach is applicable to any area where similar problems are encountered. Quantities of particulate matter reported in analysis of stack effluents indicate the need for standards of emission. Many types of dusts and fumes that contribute materially to the air-pollution problem are discharged into the atmosphere at a lower rate than 0.40 grain per cu ft at standard conditions. On the other hand, some stacks discharge at a higher

Jan 1, 1951

By V. F. Parry, E. O. Wagner

This paper summarizes investigations during 1949 on three pilot plants for drying low-rank fine coal by entrapment in hot gases. Detailed operating results on processing seven coals having moisture ranging from 24 to 62 pet, formulas and data for calculating performance of large units, and general conclusions on the problem of drying fine coal are presented. THIS paper presents a summary of investigations by the Coal Branch of the Bureau of Mines during 1949 on the removal of internal moisture from low-rank coals by processing them in the entrained and fluidized state. A paper on this subject, giving a summary of the early work, was presented at the 1949 AIME annual meeting1 and gave a the- oretical discussion of the time and heat required to dry various sizes of low-rank coal. The investigations on two pilot plants were presented in the foregoing paper. During 1949, three additional pilot plants were built to study phases of the drying problem and to investigate different techniques of handling fine coal in the entrained and fluidized state. The present paper gives operating data on the new units and presents detailed data on the drying of seven coals having bed moisture ranging from 62 to 24 pct. The problem of removing surface moisture from fine coal is principally one of heat exchange and dispersion of the coal particles so that heat can reach the liquid moisture to permit evaporation, and it does not involve the factors of size and time of contact to any great extent. If enough heat can reach the surface of the particles at temperatures higher than 250°F, moisture evaporates almost instantly, and the factors of time and size are of secondary importance compared with the factors of heat balance and dispersion of the coal. Therefore, it does not appear to be necessary to employ a fluidized bed when removing surface moisture, and the most efficient process would be one in which the coal particles are entrained in a hot gas stream of optimum temperature traveling at a velocity that causes the particles to disperse into a low-density phase. When low-rank coals are dried, the fluidized bed is necessary to provide control of time of contact. This is done by adjusting the superficial velocity which controls the density or dispersion of the particles. The density of beds having fluidized properties suitable for drying low-rank coals appears to range from about 5 up to about 15 lb per cu ft. When the density is below 5 lb per cu ft, which represents a low-density phase, optimum conditions for removal of surface moisture should be found, and this requires superficial velocities of about 15 fps. The problem of removing bed moisture from low-rank coals involves consideration of size of coal, temperature, mass flow, heat balance, pressure, time of contact, and method of dispersing the coal in the heating mediums. The experimental work presented here gives data on these factors. Coal sizes up to 1/2 in. x 0 were dried with hot gases having initial temperatures ranging from 1900° to 2400°F. Coal

Jan 1, 1951

By H. L. Teichman, E. M. Sipprelle

ENACTMENT' of Public Law 290 by the 78th Congress authorized the U. S. Department of the Interior, Bureau of Mines, to conduct an experimental program to develop the technology for obtaining oil from oil shale. In adopting and later extending this legislation, the Congress recognized the impending necessity of supplementing ground petroleum reserves with synthetic fuels. Under the provisions of this legislation, the Bureau of Mines, among other things, was charged with the responsibility of developing mining techniques, methods, and equipment for mining the oil shales of the Green River formation of Colorado, Utah, and Wyoming. The oil shales of western Colorado are apparently richer, more accessible, and more amenable to exploitation than elsewhere in the Rocky Mountain region. The site chosen for the Bureau's Experimental mine is about 10 miles west of Rifle in northwestern Colorado. It is within a 1000-sq-mile area from which, it has been estimated, 300 billion barrels of shale oil could be produced from a 500-ft measure near the top of the formation. One hundred billion barrels of this amount could be produced from the Mahogany ledge, a 60 to 100-ft section near the bottom of the 500-ft measure. This ledge is considered to have economic importance at present. The Green River formation was laid down as sediment in the bottom of vast, shallow inland lakes during Eocene time. The deposit is flat-lying, and there are no faults, fissures, or local rolls. Oil shale is actually a strong, tough magnesium marlstone, which will stand unsupported over relatively wide spans. These and other natural physical characteristics favor mechanized, low-cost mining, which is essential for establishment of an oil-shale industry. It was realized from the outset that an extensive research program would be necessary to develop mining methods, equipment, and techniques for a mechanized, low-cost operation. The program was designed to include research. into all the productive phases of mining, such as drilling, blasting, loading, transportation, and maintenance of the mine structure. The methods, equipment, and techniques developed as a result of this research have established a production of 116 tons per man-shift total labor at a direct cost of $0,292 per ton. Another important phase of the research program that has received little publicity because of its theoretical nature is study of the roofstone behavior and determination of mine structure stresses. This paper purposes to discuss this phase of the research program. Preliminary studies of the physical properties of the Green River oil-shale formation were made in the Barodyramics Laboratory at Columbia University during the latter part of 1945 and the early part of 1946.* The purpose of these studies was to determine the maximum size of unsupported underground openings that would be commensurate with safety and still permit the use of large, efficient mining equipment. Also to be determined were the pillar support to extraction ratio and the shape, size, and spacing of supporting pillars. Selected samples of possible roofstones near the top of the Mahogany ledge, as well as representative samples of different rock types found within the ledge, were obtained from the Bureau's oil-shale mine for these studies. The maximum safe unsupported roof span calculated from this work was 200 ft. Using a safety factor of four, it was theoretically determined that openings 60 ft wide could be advanced under a roofstone at the top of the Mahogany ledge. To support the overburden, 60-ft-sq pillars would be left in a checkerboard pattern. From visual observations made of core samples through the selected roofstone at the oil-shale mine, it was determined that the roofstone was actually a plate 6 to 8 ft thick. Because the calculations were theoretical and allowance had to be made for unknown cracks and fractures in the formation, openings 50 ft wide and pillars 60 ft sq were originally contemplated in the Bureau's Experimental mine. This would be the minimum allowable width that would permit use of large underground mining equipment. For lower mining costs and greater efficiency larger openings were desirable. Different but analogous approaches were made to the problem at the Bureau of Mines Applied Physics

Jan 1, 1951

By W. J. Tait, A. E. Craig, E. P. McCurdy

The Circuitron herein described applies current from the classifier motor circuit and energy from the sound of grinding media to move an oscillating disc. The disc through a photoelectric cell controls the speed of the ore feeder and consequently the amount of ore delivered to the mill. THE Telluride Mines, Inc. at Telluride, Colo., is operating a flotation concentrator recovering lead, zinc, and copper with some gold and silver. The nominal capacity is 500 tons per 24 hr. The ore, which comes from several different mines, is characteristic of the district, changing from extremely tough rhodonite to soft gouge. The ore is crushed and washed, recrushed, and delivered to the mill bin at about % in. to serve as ball mill feed. Details of Grinding Circuit The material from the fine-ore bin is carried over a belt feeder driven by a Reeves vari-speed device. The belt discharges to a No. 86 Marcy ball mill, which operates in closed circuit, with jigs and a 6-ft Dorr classifier. A Massco density controller holds the density of the classifier overflow at a predetermined point. The overflow, which contains about 5 pct + 48 mesh, goes to the flotation section. Manual Operation Before the adoption of automatic control for the grinding circuit, this work was all done manually by the operator. The bin gate was set to deliver 40 to 50 lb of ore per lineal foot of belt, and the speed of the belt was adjusted by the operator to deliver at the rate of about 500 tons per 24 hr. As the ore changed in hardness, specific gravity, or in the amount of moisture, or, as the size of the feed changed due to bin segregation, crusher setting, or other causes, the operator made corresponding changes in the rate of feed by manipulation of the Reeves control. Sometimes these changes occurred so rapidly that they caused sharp, pronounced surges, which created very considerable unbalance from normal conditions in the grinding circuit and, more critically, in the flotation section. Since it is not possible for the operator to anticipate or follow such changes closely, a very considerable unbalance from normal conditions may occur, which will require a drastic change in the rate of feed. These surges upset the fineness of grind and interfere radically with the balance between tonnage and reagent control. All of this tends to reduce the grade of concentrates, increase the loss in the tailing, and thus decrease the recovery of values. The Circuitron Some years ago, The Mine and Smelter Supply Co., in conjunction with the operators at the Telluride Mill, undertook to build a device that would give close control of the rate of feed from the ore bin and to coordinate with this a related device that would regulate the amount of sand load on the classifier rakes. Emphasis was strongly laid on the fact that such a control must apply to the entire grinding circuit and must include the feed to the ball mill, the classifier sand return, and the density of the classifier overflow. If this could be done successfully, we felt that the circuit could be operated under best possible conditions, with compensations

Jan 1, 1951

By O. C. Ralston, F. Fraas

VIRTUALLY all commercial use of electrostatic separation has employed separators depending on differences of conductance of the broken, solid mixtures treated by them. The two main types of conductance separators have been: (1) Those in which the separating field is as near a static field between two charged electrodes as it is possible to maintain in the face of the usual leakage that takes place between electrodes with some 5 to 50 kv tension between them and spacing of 1/2 to 3 in.; (2) those in which a convective discharge of corona or spray type, but not a spark discharge, takes place between the carrier electrode and a charged electrode facing it, which is of very restricted area, such as a row of needle points or a small diameter wire parallel to the face of the carrier electrode. The present paper deals with this latter type. The principle is usable in separators where the carrier electrode, usually grounded, has the shape of a roller, a conveyor belt, a vibrating table, a sloping chute, or a nearly vertical plate. The innovation studied by us has been the use of the convective field for charging material on the carrier electrode and succeeded by a field between the carrier electrode and an electrode made of a true dielectric which cannot take on a charge by conduction but is charged by being in the radius of the corona discharge. The acquired charge is of the same sign as that of the corona but is not dischargable by contact with a good conductor. However, high-tension charge tends to leak off because of the ionization of the ambient air. The field between this electrode and the carrier electrode is as near static as it is possible to obtain. There are some differences in behavior of this electrode and that of the gas tube electrode used by Sutton, as will be discussed below. Both electrodes cause increased adhesion of nonconductive particles to the carrier roll. The particles are pinned there because they are convectively charged on their outer surfaces facing the corona electrode. The electrodes also cause increased rate of release of conductive particles tending to be pinned by corona charging but conveying the charge to the grounded carrier roll almost as rapidly as it is received by the particles. The sign of the potential on the corona-emitting electrode makes little difference in the separation. An electrode usually emits negative corona somewhat more easily than it emits positive corona, as electrons are then the fundamental particles emitted. With a positive electrode the flow of electrons involved in positive corona discharge is onto the electrode from the ambient gas. Theory In the separator illustrated in fig. 1, the nonioniz-ing electrode is a rotating, nonconductive dielectric cylinder designated as B. The ions emitted by the corona electrode, or wire F, charge both the particles on the carrier electrode, or roll A, and the surface of the dielectric electrode B. Whereas the space between roll A and wire F contains ions, the space between electrode B and roll A is a pure static field free of ions. The dielectric electrode has been described by Bullock.1 However, the corona electrode, which Bullock uses for the purpose of charging only the dielectric electrode, is adapted here to also charge the particles on the carrier roll. In operation the ionic field from the corona electrode is balanced with the static field from the dielectric electrode. The ionic field charges the particle so that it adheres to the roll. The static field intensifies the discharge of the corona charge by interfacial conduction and permits the release of the particle. The ionic and static fields therefore have opposing effects.2 The values of both fields are adjusted so that the poor conductor will adhere and the good conductor will be released.

Jan 1, 1951

By S. F. Ravitz, S. R. Zimmerley

High-grade cobalt concentrates were produced from the complex Blackbird ore with very good recovery in continuous pilot-plant operations in which a low-grade bulk cobaltite-pyrite flotation concentrate was roasted in a multiple-hearth furnace to oxidize the pyrite selectively. The cobaltite was floated from the resulting calcine. THE existence of the cobalt-copper deposits in the Blackbird district, Lemhi County, Idaho, has been known for many years. However, these deposits have not yet been exploited successfully because of the low grade of the ore, absence of information on ore reserves, and lack of feasible methods for concentrating and recovering cobalt from the ore. In 1941, when it appeared that World War II might interfere seriously with imports of cobalt from overseas, the U. S. Bureau of Mines started an extensive mining and metallurgical investigation of the Blackbird deposits. Exploration by trenching and drilling1 indicated that ore reserves were adequate to warrant consideration of large-scale exploitation. A process for producing high-grade cobalt concentrates from the ore was developed at the Salt Lake City Station of the Bureau,2 and hydrometallurgical procedures for producing cobalt metal or oxide from the concentrates were developed at the Boulder City Station.3 The Howe Sound Co., through a subsidiary, the Calera Mining Co., carried out further exploratory work, which increased probable ore reserves to a tonnage commensurate with a large investment in mine, mill, transportation facilities, and refinery. Simultaneously, the metallurgical staff of the company, working in cooperation with the Bureau of Mines, further investigated the metallurgical problems. It was the opinion of both groups that pilot-plant testing was required to translate laboratory results into plant operation. In general, the Blackbird ore has a high content of pyrite and some pyrrhotite and contains 0.5 to 1 pet cobalt and 1 to 2 pet copper, which are present essentially as cobaltite and chalcopyrite. The copper can be recovered readily as a high-grade concentrate by selective notation. The cobalt then can be recovered as a bulk sulphide concentrate assaying 4 to 5 pct cobalt, but no satisfactory procedure for separating the cobalt from the pyrite and pyrrhotite directly by flotation had been found. The method developed by the Bureau for producing a high-grade cobalt concentrate is essentially an application of the old Horwood process of selective roasting.4 The bulk sulphide concentrate is roasted to form an oxide coating on the iron mineral particles, and the calcine is then treated by flotation. The oxide coating prevents the iron minerals from floating so that it is possible to obtain a high recovery of cobalt in concentrates assaying more than 20 pct cobalt. Laboratory tests by metallurgists of both the Bureau and the Howe Sound Co. had indicated the optimum roasting temperature to be 425° to 450°C. At lower temperatures the rate of oxidation of the iron minerals is too low, whereas at higher temperatures substantial quantities of cobalt are rendered water-soluble. Because of its high sulphur content (35 to 40 pct), the bulk concentrate tends to continue to burn after it has reached the optimum roasting temperature. Although excessive temperatures could be avoided in the small-scale laboratory furnaces, there was considerable doubt that the temperature could be controlled properly in industrial equipment. To provide material for the pilot-plant tests, the Calera Mining Co. installed a small, temporary flotation mill at the Blackbird property and treated

Jan 1, 1951

By L. E. Shiffman

This paper describes the operation of two kerosine flotation plants in Alabama for cleaning —10 mesh bituminous coal. One plant treats washer sludge, the other raw coal. Data on capacity efficiency and capital and operating costs is given. Methods for increasing cell capacity are discussed. IN cleaning coal it has long been recognized that methods which give excellent results for the coarser sizes may give poor or even no cleaning for the finer sizes. Effective cleaning of the fines, therefore, is usually a separate and distinct problem. For many years the problem of fines was considered of relatively minor importance. Now there is growing interest in methods of cleaning the fine sizes of coal because: 1. There are more fines. Increasing mechanization and the adoption of full seam mining methods have resulted in the production of a greater percentage of smaller particles than were produced by hand loading methods. 2. The fines have deteriorated in quality. The same causes that have led to increased quantity have also operated to decrease quality through greater contamination with slate, rock, and other diluents. 3. The fines have become more valuable. When coal was comparatively cheap, loss of the finer sizes was tolerated but, with the rapid increase in the cost of mining coal, it has become imperative that the maximum amount of clean coal be recovered from the mixture of coal and rock delivered to the tipple. 4. In many cases, the fines have become a stream pollution problem. Increasing governmental activity and legislation to prevent stream pollution often have made it necessary to retain the fine material which formerly was permitted to escape with the washer effluent into flowing streams. The Sloss-Sheffield Steel and Iron Co. operates four coal mines, all within a 20-mile radius of Birmingham, Ala., with a total annual production of 1,-250,000 tons of washed coal. Three of these mines are in the Mary Lee seam and one in the Jefferson seam. Although some steam and domestic coals are produced, the major portion of the coal is used to make coke for blast furnace operation. Fine coal is desirable for coking, and all the coal so used by Sloss is crushed to 1/4 in. Therefore, there is no problem of an outlet for even the finest sizes of coal. Since the coal is used for metallurgical purposes, low ash is of greater importance than would be the case if it were used for steam generation or domestic heating. Based on present blast furnace practice at Sloss, it is calculated that each percentage of ash in the coal carries a penalty of $0.22 per ton of coal. Thus there is a considerable margin for operating costs in a fine coal cleaning method that will result in materially lowering the ash content of the cleaned coal. Development of Fine Coal Cleaning Problem In 1942 Sloss reopened its Bessie mine and constructed at this location a new preparation plant to clean 11/4 in. x 0 coal from the Mary Lee seam.

Jan 1, 1951

By H. F. Yancey, M. R. Geer, C. L. Allyn, R. H. Eckhouse

Four coals were treated in the Humphreys spiral concentrator, and the products were examined by float-and-sink and screen-sizing tests to determine fundamental performance characteristics. The efficiency of the separation between coal and impurity was shown to be influenced greatly by particle size. FOUR coals exhibiting different washability characteristics were washed in the Humphreys spiral concentrator and the products examined by float-and-sink and screen-sizing tests to determine the mechanics of the separation between coal and impurity. With one coal, each 1/2-in. increment of the full width of the stream in the spiral was examined separately to determine where each size and specific-gravity fraction of the feed was stratified. The basic performance characteristics of the spiral were found to be the same for all of these coals. The coarsest fraction of heavy impurity stratified so far out in the stream that it could not be removed through the refuse ports and thus either entered a middling product or contaminated the washed coal in the case of a two-product separation. Impurity particles finer than about 100-mesh also were carried out in the main body of the stream and therefore were not removed in the refuse product. Little loss of clean coal in the refuse occurred in sizes coarser than 28-mesh or finer than 100-mesh, but considerable coal of intermediate size stratified in the stream that was drawn off through the refuse ports. Because of this modifying influence of particle size, the spiral is unable to make an efficient two-product separation between coal and impurity without some retreatment. Recirculation of a middling product through the same spiral as a means of obtaining more efficient operation was not attempted in this investigation. The size and specific-gravity composition of the middling product are, however, such as to render it amenable to simple hydraulic classification for recovery of coal. Retreatment of a middling product would improve the efficiency of the spiral, but even when making a substantial amount of middling, considerable clean coal enters the refuse product. Consequently, retreatment of a combined refuse-middling product would appear to offer greater promise for providing maximum efficiency. Because of the sizing characteristics of the spiral, a classified feed can be treated with higher efficiency than is possible on a natural raw coal. For this reason, semiclassified feeds, such as silt-bank or classifier-underflow materials, can doubtless be treated with higher efficiencies than those shown for raw coals in this report. Performance is, of course, only one of many factors that enter into the choice of cleaning equipment. The fact that the one company now using the spiral for cleaning coal has built a second spiral plant is ample evidence that the inherently low efficiency of this unit may be overshadowed by the low operating costs afforded by its extreme simplicity. Introduction A number of factors enter into the selection of coal-cleaning equipment, and most of them bear directly on the final cost of prepared coal. One of these factors is performance of the cleaning unit in terms of the efficiency of the separation it effects between clean coal and refuse. The importance of performance in the selection of a coal-cleaning unit varies with the difficulty of the cleaning problem. Obviously, if the character of the coal is such that only certain types of cleaning units are capable of yielding a clean product meeting market requirements, performance is of prime importance. If, on the other hand, the cleaning problem is so simple that any type of cleaner will provide coal of suitable quality, performance is important only insofar as it affects costs by determining the amount of salable coal lost in the washery refuse. With the objective of providing the coal industry with information that is helpful in the selection of

Jan 1, 1951

By H. F. Yancey, M. R. Geer, C. L. Allyn, R. H. Eckhouse

W. M. Bertholf—This is an excellent report of a well-conducted investigation, of sufficient scope to provide generally useful information. Some years ago we had occasion to test the Humphreys spiral on the middling from our table plant. The conclusions reached in this investigation come as no surprise to us. In certain applications it is virtually impossible to improve on the spiral from the monetary returns point of view even though its "efficiency" may be low. We have noted that in all successful applications of the spiral the feed is "classified," hence we were somewhat disappointed in the showing the spiral made on the Poplar Ridge coal, the only one of the four which could really be considered the equivalent of a classified feed. As was noted in the paper, however, this coal contained a considerable amount of middle and high-gravity material, apparently more than enough to make up for the theoretical advantage of small particles in these categories. A comparison of particle size in the various specific gravity classes of the four coals is given in table XXV. Table XXV. Comparison of Particle Size in Various Specific Gravity Classes. (Unit size, 150 m to 0.105 mm) Black Eagle Clements Poplar Ridge Rosly,n 1.3 Float 7.47 10.60 8.68 9.66 1.3 x 1.4 7.89 7.96 7.58 8.20 1.4x1.6 7.24 8.36 5.58 6.85 1.6 x 1.8 7.98 8.31 3.42 7.33 Sink 1.8 7.85 7.13 5.09 7.41 Avg 7.69 10.08 7.15 8.38 This tabulation indicates that there was very little difference in the size of the various gravity fractions for three of the coals, Poplar Ridge varying the most. Examination of figs. 3, 6, 7, and 8 reveals a general similarity of behavior for all coals except Black Eagle. Would the authors care to speculate on the reasons for the difference in the distribution curve for the sink 1.60 material in the washed coal of fig. 3 as compared to the others? This difference is particularly noticeable in the 8xl4-mesh size. Was there a "shape effect"? J. D. Price and W. M. Bertholf—The performance data on three of the coals studied appear to be typical, but there are certain departures from the general pattern in the case of the Poplar Ridge coal which are large enough to deserve special consideration. For example: 1. Despite the high proportion of sink 1.60 material in the feed, it was not possible to make a corresponding proportion of high-ash reject. (See table XVI.) 2. Even though approximately 70 pct of the feed was 1.40 float with an ash content of 6.6 pct, and this

Jan 1, 1951

By J. D. Clark

ONE'S lifetime does not usually offer the oppor-tunity to observe and be a part of the commercial development of an entirely new material. Petalite has been known for over a century, but at first it was presumed rare. However, it was found to be so abundant that it has become a new and extremely useful industrial tool. Mineralogy Petalite is a lithium-aluminum-silicate mineral. It was from petalite that J. A. Arfwedson first discovered lithium in 1818. The theoretical composition of Li2O • A12O8 • 8SiO2 is equivalent to 78.4 pct silica, 16.7 pct alumina and 4.9 pct lithia. The mineral in its known commercial occurrence is colorless to white and rarely gray, transluscent and rarely transparent. It is brittle, has a vitreous luster, and occurs usually in cleavable masses (fig. 1). It is classified as a di-silicate, but is similar in composition to the metasilicate spodumene, Li2O.-Al2O8-4SiO2 (table I). However, their specific gravities are quite different (3.13 to 3.20 for spodumene, 2.39 to 2.46 for petalite). Occurrence Petalite specimens have been reported in many localities; however, there are only three known countries having occurrences of commercial significance. The classic locality is Sweden's Utö Mines, which, although reportedly still containing considerable petalite, have been closed and are not expected to be operated in the near future. At Varuträsk, in northern Sweden, petalite carrying 4 pct lithium oxide is the main ore. In 1946 there was reported about 1000 tons on hand. However, the iron content was 0.60 pet too high for other than chemical use. Petalite has been reported in western Australia. At the Cottesloe mine, where small quantities have been reported available, it is produced in conjunction with feldspar. Petalite occurs in considerable quantity in the feldspar quarry at Londonderry, and it was estimated1 in 1943 that over 3700 tons might be available. The petalite in South West Africa was first discovered by F. J. Jooste, a lawyer by profession, in 1938. The following year he discovered a much larger deposit, and in 1948 another important discovery was made. These deposits occur near Kari-bib and constitute the only known commercial deposits. Their full extent has not yet been determined, but huge massive blocks of petalite of the purest grade, running into many thousands of tons, can be seen in all these deposits. To date, 60,000 tons of high grade, low iron containing ore has been blocked out in the major bodies. Mining for years to come will be confined to simple quarrying. Selection from accessory minerals has proved

Jan 1, 1951

By R. Gibson, F. S. Turneaure

The tin deposit of Monserrat, Bolivia, consists of one major vein 1600 m in length. The ore is unusual because of the notable quantity of teallite, even though cassiterite is the principal tin mineral. The deposit, formed at shallow depth under a wide range in temperature, may be classed as xenothermal. Polished sections reveal a complex history of replacement, with low-temperature minerals deposited before high-temperature minerals. THE Monserrat mine1'2 of the Compania Minera Monserrat is located 11 km east of Callipampa, a station on the Antofagasta and Bolivia Railway some 55 km south of Oruro. The mine lies in a group of low hills on the east slope of a broad valley of northerly trend. The valley is separated from the altiplano to the west by a prominent ridge that rises several hundred meters above the surrounding country (fig. 1). Callipampa is situated a short distance west of the ridge, near the east margin of the altiplano. The elevation at Callipampa is 3700 m and at the mine 4100 m. Monserrat is near the northeast limit of what may be called the Poopo-Pazna district, an area of northwesterly trend about 25 km in length and 15 km in width. The district includes the tin-silver prospects of Poopo to the north, the zinc-tin deposits of Salvador to the southwest, and the tin-tourmaline veins of Avicaya to the south. Along the prominent ridge already mentioned, which forms the western limit of the district, there are several tin and antimony prospects including the Trinacria mine, which is famous for its fine specimens of cylindrite. Although the principal ore mineral is cassiterite, the tin deposit of Monserrat contains an unusual amount of teallite. Cassiterite occurs in part as a finely granular mixture with sulphides of iron and zinc and in part as needle tin of late hypogene age. The regional structure is dominated by broad folds typical of the Andes of Central Bolivia and similar to those of the Llallagua district. The area adjacent to the Monserrat mine is underlain by thin-bedded shale somewhat variable in color but predominantly dark grey to black. The beds strike northwest and dip at low angles southwest, forming part of the northeast limb of a major syncline. A resistant sandstone formation, which underlies the black shale, crops out along a ridge east of the mine and also accounts for the high ridge to the west that was previously mentioned. No igneous rock of any kind is found at Monserrat. The nearest intrusive bodies known are the irregular dikes and masses of quartz porphyry at Avicaya about 10 km to the south. Mine Workings The Monserrat property includes one main vein, which can be followed on the surface for 1600 m, and two or three branch veins of distinctly minor importance. The principal haulage tunnel, referred to as the San Carlos adit or level III, extends northeasterly for 1700 m, following the main vein throughout most of this distance. Above the adit, there are

Jan 1, 1951