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THE BULOLO gold-dredging operations were described at some length by Frank W. Griffin at the Oct. 11 meeting of the San Francisco Section. This enterprise is on the island of New Guinea, about 1800 miles north of Sydney, Australia. The area in which the present placers exist originally was German territory but later the location was taken over by Australians. Gold was discovered in 1923 by prospectors who required 8 to 9 days to pack into the country over severe mountain trails. In 1926 a rich discovery was made on Edie Creek and in 1927 the deposit on the Bulolo River was discovered.

Jan 1, 1932

Located 19 km (12 miles) north of the Carlin gold mine and 43 km (27 miles) north of Carlin, Nevada, is Carlin Gold Mining Company's Bootstrap mine. This deposit is similar in character to the Carlin ore, except it occurs in the upper plate formation of the Roberts thrust fault rather than in the lower plate rocks. The deposit was discovered over 50 years ago and operated on a small scale in the early 1960s using vat cyanide leaching. Newmont Mining Corporation acquired the property in 1967, and from an open pit mined 650,000 mt (717,000 st) of ore which was trucked to the Carlin conventional cyanide plant. (See Gold and Silver Cyanidation Plant Practice by F. W. McQuiston, Jr. and R. S. Shoemaker, Society of Mining Engineers, AIME, 1975. ) With the same grade of ore, the gold recovery by conventional methods is somewhat lower than that of the Carlin ore, even with a finer grind. Leaching tests were first conducted by the United States Bureau of Mines in 1968; and in 1971, ‘72, Carlin conducted extensive cyanidation investigations by bottle testing, simulated heap leaching in tall columns, and small scale pad heap leaches. Gold extractions ranged from 55% to 95% with low reagent consumptions. In 1974, a 545 mt (600 st) heap leach test was made at Carlin on pit-run low-grade material assaying 1.37 g per mt (0.04 oz per st). In a six week period the extraction of gold was 47.57'0. Carlin uses atomic absorption analyses on cyanide leach liquors to determine potential leach ore grade rather than fire assaying because it gives a more practical assessment of leach recovery than does fire assaying, which accounts for the total gold content. Before open-pit mining was started, plans were made to heap leach, the ore at Bootstrap which was too low grade for transport to Carlin. The topsoil outside of the designed pit area overlaid ancient lakebed clays which were found suitable for compaction to form an impervious base on which the dumps to be leached could be placed. By using the cost of truck haulage and milling at the Carlin plant and the price of gold 'at that time, a cutoff grade was established which resulted in the gold content of the first 150,000 mt (165,000 st) of the heap leaching ore being 1.51 g per mt (0.044 oz per st). An additional 500,000 mt (552,000 st) of lower grade material, assaying 0.96 g per mt (0.028 oz per st), is available for leaching as economics permit.

Jan 1, 1981

By F. Wartenweiler

Much has been written on milling and cyanidation as practiced on this gold field, therefore the purpose of this paper is to record its development only briefly and to describe in more detail the contemporary state of the science and art of this branch of metallurgy. It is a pleasing task, because a steady evolution is to be recorded. Definite trends influenced the development. Some have the stamp of original thought and observation; many have been achieved from a logical sequence of planned experimentation and research; while others have resulted from the drive of sheer necessity. The large scale of operations and capital outlay has compelled conservative progress. Banket ORe For a better understanding, a review of the characteristics of the ore may not be out of place. The Witwatersrand gold ore is named "banket" from the similarity of the oxidized outcrop ore to the Boer almond cake of that name. The inter bedded orebody is commonly called "reef." It is a conglomerate composed of various sizes of pebbles, principally non-gold-bearing white quartz, which constitute about 70 per cent of the bulk, within a matrix of cementing material consisting of fine-grained quartz, sericite, chlorite and chloritoid and containing the precious metals and mineral sulfides. Argillaceous components in the mil pulp may be as high as 32 per cent. Of the sulfide minerals, pyrite and pyrrhotite are predominant. Most base-metal sulfides are present, but sporadically and generally only found in highly concentrated mill products. In the latter, osmiridium is recovered in varying amounts; from 1 oz. in 3000 tons to 1 oz. in 10,000 tons of ore. Most of the oxidized ore has been mined, so milling is confined principally to the sulfide ore. A typical analysis of sulfide ore is: silica (SiO2), 86.76 per cent; pyritc (FeS2), 2.75; ferric oxide (Fe2o3), 2.65; alumina (Al2o3), 6.91; lime (CaO), tr.; magnesia (MgO), 0.70; sp. gr., 2.7. Milling grade at present price of gold (137 shillings per ounce) varies from 3.5 to 8.5 dwt., with one mine milling ore as high as 16 dwt. The

Jan 1, 1935

By R. E. Barnes

An overall view of the perlite industry is concisely presented. The geology, mining, milling, processing, and applications of perlite, as well as the present status of the perlite industry are treated. Perlite, in strict petrological usage, describes a specific variety of volcanic glass in which strain, incident to cooling, yielded a concentric structure of fracturing which may be visible either to the naked eye or under a microscope. In commercial usage, the term perlite includes any naturally occurring acidic glass of volcanic origin that will expand when heated to a suitable temperature. The name perlite is a derivation of perlstein which originally defined "certain glassy rocks (hyaloliparites, hyalo-rhyolites) with numerous concentric cracks, from the fancied resemblance of broken out fragments to pearls."1 While perlite has been known to geologists for many years, it was not until 1941, while investigating perlite for use in enamel, that an assayer in Superior, Ariz., noted the unusual expansion characteristics of the ore. Further investigations were delayed until following World War II when many small plants were built and experimentation in the treatment and uses of perlite began. The Perlite Inst. was organized in 1949 to promote the development of the new industry, establish commercial standards through research, and to explore new uses and markets. The Institute, located in New York, represents some 50 producers in the U.S., Canada, Australia, England, France, Germany, Greece, Japan, Mexico, and New Zealand. In 1949, 71,l002 short tons of crude perlite were produced, sold or used by U. S. producers, and this figure increased to 325,0003 short tons by 1959. Such a rate of increase indicates the rapid and continuing growth of the perlite industry. GEOLOGY Perlite ore in its crude form has been found in seven of the mountainous western states, since un-devitrified siliceous volcanic rocks are limited to this section of the U. S. Commercially operated perlite deposits are normally of Eocene and Oli-gocene age.4 These deposits are often several hundred feet thick and may extend over hundreds of acres. In 1959, 13 companies in six states produced crude perlite, although the output in New Mexico comprised 79 pct of the total domestic crude output. Other states in order of their production were: Nevada, Arizona, California, Colorado, and Utah.5 Foreign deposits of perlite which are either producing or under exploration are Australia, Canada, Greece, Hungary,6 Iceland, Ireland, Japan, Mexico, New Zealand, and Sardinia. The ore may exhibit a variety of characteristics depending partially on its water content and eruption history and may range through many shades from black to white. While these characteristics may vary slightly among deposits, perlite ore from a particular source is normally quite uniform. Available chemical analyses of perlites from a large sampling of domestic deposits were recently assembled by the Technical Committee of Perlite Institute and are presented in Table I. While the ranges given do not apply to a particular perlite, they are an indication of the chemical analysis which may be expected from any presently commercially operated deposit. The specific gravity of the perlite ore ranges between 2.2 and 2.4, and its bulk density when crushed and sized is 75 to 90 Ib per cu ft. MINING AND MILLING Perlite ore is quarried from open pits by conventional power equipment, although blasting has been occasionally employed. The rock is then stock-piled until needed. While the milling of perlite may seem a simple process compared to that required for basic metals, or even other nonmetals, the need for accurate control of particle size becomes extremely important in further processing. The first step in this operation is the primary crusher, which is normally a jaw crusher. Following crushing, the material is dried, if necessary, in a rotary dryer and conveyed to a secondary grinder which may be a rod mill, hammer-mill, or cone crusher. However, the crushing characteristics of the perlite, caused by incipient cooling fractures, requires it be removed from the grinding operation once it is reduced to the required size so as to prevent overgrinding. Following the grinding operations, various size screens are used to separate the ore according to particle sizes required for particular uses. Most plants use two or more stages of crushing with perhaps four stages of screening.

Jan 1, 1961

By E. W. Greene, J. B. Duke

This paper deals with the application of froth flotation techniques to the beneficiation of kaolin clay and phosphate slimes, two very fine particle sized materials. The kaolin problem involves the removal of a small amount of a contaminating mineral (anatase), while the phosphate problem involves the concentration and recovery of a much larger percentage of apatite. Due to these differences, and also different response to flotation, the two problems are discussed in separate parts of the article. An idea of what is meant by "Ultrafine Minerals Or Slimes" as used in the title can be had by an examination of Fig. 1. Particle size distributions are presented for a crude kaolin clay, a coarse fraction and a fine fraction. The crude kaolin as mined was blunged with water, degritted through a 325-mesh Tyler screen, and dispersed with sodium silicate. The dispersed crude was then centrifuged in the laboratory into the coarse and fine fractions shown in the figure. The distribution curves were obtained by the well known sedimentation technique of Casa-grande1 using a long arm centrifuge to obtain equiva- lent spherical diameters down to 0.2 p on the finer fraction. Fig. 1 shows the coarse fraction kaolin to be 50% finer than 3 p, while the crude kaolin is 509 f'lner than 1.4 p, and the fine fraction 50% finer than 0.46~. For comparison 53 p on the horizontal scale would represent Stokesian diameters which would just pass a 325-mesh Tyler screen, taking into consideration the Andreasen sieve correction factor of 1.22 times the screen opening. 2 Another helpful comparison is the surface areas, which in square meters per gram, are 0.044 for 53 p particles, 0.77 for 3p, 1.65 for 1.4 p and 5.0 for 0.46 p particles. Fig. 2 is an electron micrograph of a similar fine fraction kaolin showing the hexagonal plate-like particle shapes.

Jan 1, 1962

By R. C. Neavel

One of the most important requirements to be met by any coal producer is delivery of size consists tailored to specific applications. Generally these specifications include a minimum as well as a maximum size limit. The customers of a major Indiana coal producer noticed that some shipments of sized coal contained a large proportion of undersize pieces. Also, in these shipments, the pieces which remained within the specified limits were cracked and, when handled, crumbled readily. The producer, certain that these shipments had been properly sized at the tipple, suspected that degradation had occurred in transit. The Indiana Geological Survey was requested to examine the coal from one of these shipments to attempt to determine the cause of breakage. Two basic facts about coal and coal seams led to the employment of petrographic techniques in this investigation. First, all coal seams consist of accumulations of diverse organic and inorganic constituents; and second, most coal breakage, either desirable or undesirable, can be attributed to the inherent weakness of some of these constituents or to poor adhesion between certain constituents. Analyses of degraded coal Sampling: Samples were secured from 2 stock piles of 2-by 4-inch coal in a dealer's yard. The coal in both piles had received the same treatment from mining to storage, although it had been mined and shipped on two different days. The size consist of one pile, however, appeared smaller than that of the other. The sample from the degraded pile is referred to as the slack sample, and the other is called the unslack sample. Size analyses: In order to determine the actual size consist, the samples were screened, with a minimum of rough handling, on square 2-inch, 1-inch, 1/2-inch, and 12-mesh screens. It is assumed that 100 percent of the 2-by 4-inch coal is retained on a 2-inch round screen at the tipple. Analyses of several loads at the tipple indicate that approximately 26 percent of the 2- by 4-inch coal, 100 percent of which was assumed held on a 2-inch round screen, passed through a 2-inch square screen. Twenty-six percent may be taken as the normal difference between round and square screens. Results of the analyses of stockpile coal are shown in Table 1. Cumulative percent versus size are plotted on a Rosin-Rammler chart as Figure 1. Both samples approximately follow normal size distribution; this indicates that breakage is not size preferential and also that the samples are representative of the stockpiles. Although both samples exhibit the same general size distribution, the median size of the slack sample is considerably smaller than that of the unslack sample. Fifty-five percent of the unslack sample was undersize (29 percent greater than the normal 26 percent difference), and 83 percent of the slack sample was undersize (57 percent greater than the normal 26 percent difference). Both piles contained a greater percentage than normal of undersize coal; this indicates that they were both affected by degradation, but to a different degree. Bedding-plane analyses: Coal, because of its laminated nature, will break preferentially along bedding planes, especially where adjacent constituents are poorly cemented to each other. Subsequent breakage perpendicular to the

Jan 1, 1961

By F. E. Jaumot, A. Sawatzky

IT as long been recognized that there is a need for accurate systematic data on diffusion in solids. Originally, such data were needed to check existing theories but, more recently, it would appear that it could be used most advantageously as a guide to, or a check for, much needed extensions of the theory of diffusion. The awareness of the conditions which are important to experimental accuracy, the availability of the materials for accurate experiments, and the techniques have improved to the extent that it would now seem to an experimentalist that, unless extensions of the theory are made which take into account the fundamental parameters of the atoms, the electrons, and the lattice in some detail, the field of diffusion will rapidly become an empirical or semi-empirical field. The present series of experiments was designed to provide accurate data on the diffusion of the elements of the IB and IIB subgroups in silver; that is, to provide data on the diffusion in silver of three face-centered-cubic elements with roughly similar chemical properties, but with different atomic or ionic sizes and, similarly, for three noncubic elements. The elements of these two subgroups have different valences, so that the data will not give any significant information on differences due to the crystal structure of the solute. The present paper reports data on the diffusion of copper and mercury in single crystals of silver; the data for the other elements of the IB and IIB subgroups were reported previously by the authors1'2 and by Tomizuka and Slifkin.8 Experiment The general technique used was identical to that previously described1,2 and will not be repeated here. In all cases, the sectioning technique and high specific activity isotopes" were used. The short half-life of Cu64 made decay corrections more important than for the other isotopes, and necessitated around the clock shifts while data were being taken. Each section was counted at least twice (as was true for all longer half-life isotopes) and a monitor was counted along with each sample. Data for the diffusion of mercury in silver are presented at a greater number of temperatures than for the other solutes. This is a result of two circumstances. First, six (of 30) samples had to be discarded because of pitting or recrystallization which occurred during the diffusion runs, even though the samples had sustained high temperature prediffusion anneals. Either of these was a very rare phenomenon in the other experiments, and their occurrence in this case is considered in more detail below. Second, the curves of the logarithm of the activity of the sections vs the square of the penetration deptht were, in general, quite good, but frequently seemed to show one point to have high activity and the next point to have slightly low activity in terms of the best straight line. When this occurred, there was remarkable regularity of the high-low sequence. In any event, the results, as far as diffusion coefficients are concerned, were not affected by this amalgam effect. In fact, the mercury data had the highest internal consistency of any of the data. Results Diffusion of Copper in Silver—-Diffusion runs were made at eight temperatures In the temperature range from 700° to 950°C, for times ranging from 5 to 45 hr. A linear curve of the logarithm or the activity vs the square of the penetration depth was obtained for all samples. Table I summarizes the results for individual samples and Fig. 1 gives a plot 02 the logarithm of the diffusion coefficients as

Jan 1, 1958

By Robert Linton

The Sierra Consolidated Mines Co., organized in 1909, owns, together with other holdings, practically all of the productive mineral area in the Ocampo district. Lying within this area are 15 mines, large and small, that hid previously been worked and appear to have yielded, since the discovery of the camp, about 100 years ago, bullion of a total value of over $50,000,000. The deepest mine, the Santa Juliana, had been worked to 1,100 ft. below the outcrop. The reopening of the mines involved, first, extensive and thorough development to open up the known ore shoots at greater depth and explore the ledges for additional shoots, and, second, a revision of existing equipment and methods, so as to increase output and reduce cost of production. This paper will give some account of the metallurgical treatment and a description of the new mill to be built at El Salto. The Ore The ore is practically all argentite, Ag2S, with some associated gold, and carried in a siliceous gangue. The typical ledge matter is an ande-sitic breccia cemented with quartz, or quartz stockwork in andesite.l The proportion by weight of silver to gold averages about 60 to 1. The silver sulphides occur ordinarily in rather fine. dissemination, sometimes not being easily discernible with the naked eye, even in ores of very good grade. The ledges also contain considerable pyrites, which as a rule is barren. "Bronce" is the local term applied to such barren pyrites, and "Pasta" to such as carries values.

Jan 1, 1915

By R. W. Hertzberg, F. D. Lemkey, J. A. Ford

The effect oj unidirectional solidification upon the microstructural, crystallographic and mechanical characteristics of high -purity A1-A13Ni eutectic a1loy specimens has been investigated. varying the growth rate caused the morphology of the Al3Ni phase to gradually change from platelets to rods. In addition, the effect of increasing the rate of planar liquid-solid interlace movement is to decrease the Al3Ni rod diameter and inter-rod spacing. The platelets and rods grow approxinlately parallel to each other and are aligned in a faulted substructure. A single unique crystallographic relationship hetween the A1 and ASNi phases was found and may he descrihed as : interface {001} Al3Ni {331} Al and growth direction <010>Al3Ni || <110> Al The platelet interlace and the aligned rows of rods are both uniquely defined by the above statements. Alignment of the Al3Ni phase by unidirectional solidification has given rise to a threefold increase in strength over that exhibited by specimens with an as-cast microstructure. These results illustrate the possible use of this eutectic alloy as a zohisker -reinforced structure. It has been demonstrated by Winegard et al.,1 Kraft and Albrighht,2 Chilton and winegard,3 Chad-wick,4 Yue,5 Tiller,6 and others that a planar liquid-solid interface may be established in binary eutectic alloys by proper control of heat flow during the solidification process. The unidirectional movement of such an interface results in a eutectic mi-crostructure consisting of an essentially parallel array of the two phases over an entire ingot. Two dominant phase micromorphologies have been produced using this technique: that of substantially parallel alternating lamellae of each phase or long thin parallel rods of one phase imbedded in a continuous matrix of the other phase. While not all eutectics can be controlled in this manner, it has been shown that a "normal" eutectic may, under properly controlled conditions of purity, liquid and solid thermal gradients, and solidification rate, be forced to solidify with essentially parallel phase particles. Kraft and Albright2 have demonstrated that, when the growth rate is too rapid or when the thermal gradient in the liquid at the growing interface is too low, a layer of constitutionally supercooled liquid, formed by impurity build-up, will stabilize a cellular rather than a planar interface. The unidirectional movement of the cellular interface forms the macrostructure of eutectic colonies.7 Chilton,3 Chadwick,4 and Tiller,&apos; studying eutectics of zone-refined Pb-Sn, A1-CuAl2, Zn-Sn, and Cd-Zn systems, noted that in the absence of impurities a planar interface is stabilized even when rapidly solidified. Although it is generally agreed that impurities break down a planar interface to form the eutectic-colony macrostructure, at present the origin of the varying micromorphologies (i.c., plates or rods) in a given normal eutectic alloy is not completely understood. Tiller8 has predicted that the micromor-phology produced may be dependent on solidification rate (i.e., at fast rates a rodlike structure is preferred whereas at slow rates a lamellar structure should form), and also suggested that rod formation may be favored at large phase-volume ratios. yue5 has experimentally verified Tiller&apos;s prediction in the eutectic Mg-Mg17Al12 by observing a lamellae-to-rod transition with increasing growth rate where the phase-volume ratio was approximately 2.3:1. However, Hunt and chilton9 more recently unidirec-tionally solidified over a wide range of growth velocities six different eutectic systems with phase-volume ratios between 12:l and 2.7:1 and observed no lamellae-to-rod transition. chadwickl0 has proposed that the change in micromorphology in some eutectic alloys is due entirely to the presence of impurities and further states that the lamellar structure is the characteristic structure of pure eutectic alloys even when the phase-volume ratio is as great as 12:l. Kraft11,12 has shown that the parallel lamellae in certain eutectic alloys assume a unique crystallographic relationship during unidirectional growth. This preferred crystallography may be developed

Jan 1, 1965

By Eugene A. Stephenson

NUMBER of noteworthy events in the petroleum industry may be reported for 1941, of which the most spectacular was doubtless the rise in the daily rate of crude-oil production to a peak of approximately 4,300,000 barrels. This enormous daily rate did not ensue from the furor of flush pools but represented for the most part merely a slight additional daily delivery from many wells that had previously been operated on a more restricted basis, in line with good conservation practice and also from the adroit increase in the number of days per week by the Texas Railroad Commission. Another important fact was the increase in price throughout the Mid-Continent area from the static level of $1.02 for 36°, A.P.I., crude, to $1.17 per barrel. Efforts to elevate the price above this point have met with severe opposition from the O.P.A.. as has also a similar effort by the PennsyIvania Crude Oil Producers&apos; Association, but the vastly increased costs for materials and labor may shortly give the industry the enhanced price which it has long deserved.

Jan 1, 1942

By J. Th. Erb

Oil is found in commercial quantities in the islands of Sumatra, Java, Borneo, Tarakan and Ceram. The oil deposits of the Dutch East Indies and Sarawak are with one exception of Tertiary origin and as a rule are bound to anticlinal structures. The number of layers varies in most fields and in some of them asmuch as 14 oil-producing levels have been established. The crude oil varies rather much in composition. Some of it, particularly from North and South Sumatra, has a gasoline content up to 50 Per cent. Borneo produces mainly wax oil, while the Tarakan oil can be used directly as liquid fuel. In Java both wax and wax-free oil are produced. &apos; In all important fields installations have been erected to treat the wet gas. The lean gas is used for gas-lift purposes, Marietta process, and as fuel for boilers and motors, either in the fields or in the refineries. Formerly only percussion systems (Canadian, Californian and a drilling system devised especially for the purposes of the Royal Dutch Petroleum Concern) were applied in the Dutch East Indies, but in recent years the rotary system has been predominant and good results have been obtained with it. The deepest producer has recently been struck in South Sumatra at a depth of 1624 m. (5328 ft.), with a daily production of about 350 cu. m. crude oil. The crude is brought by an extensive system of trunk lines to the refineries, of which one is in North Sumatra (Pangkalan Brandan), two arc in South Sumatra near the town of Palembang, two in East ,Java (Tjepoe and Wonokromo), one in East Borneo (Balik Papan) and one in Sarawak (Miri). Production The principal producing concerns in the &apos;East Indies are: De Bataafsche Petroleum Maatschappij, the producing subsidiary of the Royal Dutch Petroleum Concern; Sarawak Oilfields Ltd., one of the

Jan 1, 1929

By R. H. Richards, A. E. Woodward

In Rittinger&apos;s Aufbereitungskunde of 1867 occurs (p. 195) the following table, which shows the rate of falling in water of fragments of minerals, of irregular shapes and of five different specific gravities: Diameter of particles in millimeters. 10 8 6 4 3 2 1 .5 Sp. Or. Velocity in meters per second. Auriferous silver, 16..9 .8 .7 .57 .5 .4 .29 .20 Galena,. ..7.5 .62 .55 .48 .39 54 .28 .20 .14 Pyrites,. ..5.0 .49 .43 .38 .30 .26 .22 .15 .10 Quartz,. ..2.6 -30 .27 24 .19 .17 .14 .10 .07 Coal,. ..1.3 .13 .12 .10 .08 .07 .06 .04 .03 With the purpose of extending this valuable table, to cover a larger number of minerals and sizes of each mineral, and of making it more easy of application in deciding upon a sieve-scale for any group of minerals, we have prepared the curves of velocity of fall of bodies in water shown in Plates I., XI., 111. and IV., using Rittinger&apos;s formula V= 2.44d(8 —1) where V= velocity in millimeters per second. " D = diameter of particles in meters. " 8 = specific gravity of the mineral. These curves embrace a large number of specific gravities, and for each specific gravity the relation of the diameter of the mineral particle to the velocity of its fall in water is given for all diameters and velocities used in practice. The second diagram is an enlargement of the lower left-hand corner of the first, the third bears the same relation to the second, and in like manner the fourth to the third. Plate I, therefore, covers the work of the coarsest jigging 2 inches in diameter; and under

Jan 1, 1890

By Richard W. Smith

There are three distinct varieties of phosphate rock, in Tennessee, known commercially as: (a) the "brown" rock, which is the residual product of the weathering and natural concentration of certain phosphatic Ordovician limestones; (b) the "blue" rock, which is an unaltered phosphatic stratum of Mississippian or Devonian age; and (c) the "white" rock, which is the result of chemical replacement and deposition. Before 1894, South Carolina and Florida had supplied most of the phosphate of the United States, South Carolina leading until that date and Florida from then until the present time. In December, 1893, the blue phosphate of Tennessee was discovered almost simultaneously on Swan Creek, in Lewis County, by two independent parties—by one as the result of a careful and systematic search for a workable deposit of the long known kidney phosphate, and by the other through mistaking it for a "bloom" of coal. Leases were taken and, with further prospecting disclosing other deposits in the region, the development progressed as rapidly as could be expected without railroad transportation. In January, 1896, Judge S. Q. Weatherly, who was interested in the development of the blue phosphate, while riding along the road south of Mt. Pleasant, noticed some slabs of a brown rock in place above the limestone. He took samples, and had them analyzed, thus discovering the high-grade brown phosphate of this district. This discovery was kept a secret until July, 1896, when mining was started at the village of Mt. Pleasant. Development proceeded rapidly, many owners obtaining their capital by the sale of their rock fences, which were largely composed of slabs of high-grade brown phosphate. The development of the brown rock, which is richer and more easily mined than the blue, caused Tennessee to advance from a poor third in production to second, surpassing the output of South Carolina, and being excelled only by Florida.

Jan 1, 1927

By Hertha R. Freche

The diffusion of magnesium and silicon from the core into the high-purity coating of Alclad sheet is important commercially, and led to the use of a duplex product for the study of diffusion by means of spectro-graphic analysis of successive layers. It has been found that in the duplex materials the diffusion of magncsium from an undersaturated aluminum-magnesium alloy into its high-purity aluminum coating is approximately of the same order of magnitude as the diffusion of silicon from a saturated aluminum-silicon alloy into its high-purity aluminum coating. However, if magnesium and silicon are both present in the alloy in the ratio of Mg2Si, they appear to diffuse in the approximate ratio of this compound. Finally, if the alloy contains silicon in excess of the amount required by magnesium silicide, the amount of diffused silicon in the high-purity aluminum coating will also be in excess. The interdiffusion of metals in the solid state is of both scientific and technical interest, because it is helpful in the study of the alloy structure and in the solution of problems pertaining to the strength and surface condition of the fabricated product. W. C. Roberts-Austenlt carried out the first quantitative determination of the coefficient of diffusion of gold into lead. This classical experiment was followed by a series of investigations on the diffusion of metals and nonmetals in iron and steel and in nonferrous metals and alloys. Recently the diffusion of copper in coated strong aluminum alloys was investigated by A. Burkhardt2. This appeared to be the first published investigation on the diffusion of metals into aluminum. The present investigation was the direct result of a technical question in connection with alloy fabrication and is a summary of some preliminary work pertaining to the diffusion of magnesium and silicon into aluminum. The results necessarily will be subject to some variations, since this is the first attempt to evaluate the coefficients of diffusion of magnesium and silicon into aluminum according to this method of procedure. This

Jan 1, 1936

By Clarence Zener

IT is customary to regard the stress-strain relation as consisting of two parts, the elastic region and the plastic region. The essential attribute of the plastic region is the presence of a permanent set, which remains upon removal of all stresses. The essential attribute of the elastic region is the absence of such a permanent set. Absence of a permanent set does not imply a linear relation between stress and strain, nor even a single valued relation. In fact, in no real metal is stress a single valued function of strain in the elastic region. The property of a solid in virtue of which stress and strain are not uniquely related in the elastic range will be called "anelasticity." The purpose of the present article is to review and correlate the past work upon the anelasticity of metals. The following paper by W. A. West124 describes investigations upon anelastic properties of iron. Many manifestations of anelasticity have been observed, the first as early as 1835 by Weber.1,2 In his experiments Weber subjected a specimen to a constant load (or deformation) for a given time, and then suddenly removed the load. He observed that most of the deformation suddenly disappeared, but that there remained a residual deformation, which disappeared only gradually with time, as is illustrated in Fig. I. The occurrence of a residual deformation which only gradually disappeared, Weber called the elastic after-effect (Elastische Nackwirkung). An extensive review of the early work upon elastic after-effects has been given by Auerbach.3 Another manifestation of anelasticity is the dependence of the elastic constants-e.g., Young&apos;s modulus upon the method of measurement. Thus, if dynamic methods are used, the measured value will depend upon the frequency of oscillation. Still another manifestation is dissipation of energy attending vibration, called damping or internal friction. One objective of metallurgical science is the understanding of the plastic and fracture properties of metals in terms of their microstructure. Through such an understanding will come the ability to design microstructure to give the desired mechanical properties. A study of an elasticity will furnish an invaluable aid in the attainment of this objective. Such studies, when properly guided, yield information regarding the relation between stress and strain in the individual components of the microstructure. In order, however, that this type of information may be obtained, it is necessary that the experimental studies of anelasticity be interpreted by the appropriate mathematical analysis. This article, therefore, starts with the mathematical formulation of anelasticity.

Jan 1, 1946

By R. H. Adams, A. M. Khan

The Huntington Beach oil field is on the Pacific Coast, southeast of the City of Los Angeles (Fig. 1). It lies both onshore and offshore. Most of the onshore section is operated by Standard Oil Co. of California, WOI. The Tar reservoir is of upper Pliocene age, about 200 ft thick, and consists principally of fine-grained, unconsolidated, somewhat shaly sand, with thin inter-bedded shale stringers throughout (Fig. 2). The reservoir is a broad, highly faulted, elongated anticline, with an areal extent of about 370 acres. Most of the faults bisecting the structure are normal, trend north-south, and are the main cause of reservoir discontinuity (Figs. 3 and 4). From 1964 to Dec., 1967, 259 steam stimulations were performed on a total of 125 Tar sand wells. Eighty-six new wells were drilled and steam stimulated. Of the total, 96 have been stimulated twice, 34 three times and 3 four times. During this same period, a two-injector steam drive pilot has been in progress. The response obtained from this project has been compared with the response to cyclic stimulation. This paper presents an analysis of the field observations for the two methods of thermal recovery. History of Development Drilling to the Tar pool commenced in 1924 with an initial well production of 40 to 50 B/D of 12o API gravity oil. Complete primary development of the pool has been spread over several years. Approxi- mately 40 wells were drilled during the first phase of development, which was completed in 1945. The subsequent primary development started in 1960 with initial well productions of 30 to 40 B/D oil. Thirty-seven wells were completed in the Tar pool during this phase. For the most part, slotted, gravel flow packed liners have been used in the wells in this pool. A cyclic steam stimulation pilot was initiated in March, 1964. Based on the very successful stimulation in the pilot area, the cyclic stimulation project was expanded to include all of the Tar pool wells producing at the end of 1965 (Fig. 5). Also, a development drilling program was begun at this time. To date, 86 new wells have been drilled under this program. The number of wells drilled each year in this secondary phase of development is also shown in Fig. 5. New Well Completion Practices Most of the new wells have been completed to produce from both the Upper and Lower Tar zones. Generally, a casing stub is cemented from the top of Lower Tar to the base of clean Upper Tar sands to exclude intermediate water-bearing sands. The water string is then cemented from the top of Upper Tar to about 1,500 ft with neat cement and 30 percent silica flour. The annulus from 1,500 ft to surface is then filled with a petroleum-based gel through a high-temperature port collar. Once again, a history of sand-free production dictated the use of 60-mesh gravel flow packed liners. The Upper and Lower Tar

Jan 1, 1970

By Howard N. Eavenbon

The low-volatile, or smokeless, coal field of southern West Virginia is in Fayette, Raleigh, Wyoming, Mercer, Summers and McDowell counties, in the extreme southern portion of the state, and extends into the adjacent counties of Tazewell and Buchanan in Virginia. McDowell County is the largest producer, with Fayette and Raleigh counties next. The total area of the field (Fig. 1) is approximately 1600 square miles, or 1,024,000 acres. The percentage of volatile matter in the various seams ranges from 15 to 26 per cent, The upper limit of true low-volatile coal is usually considered to be 23 per cent volatile matter; above this point, the coal usually is classed as medium volatile. This limit will be used in this paper. Geological Position of Seams The smokeless seams are among the oldest coal seams in the country, as the Pocahontas group of rocks is the lowest member of the coal-bearing series and rests on top of the Mauch Chunk red shale, in or below which no coal seams of any importance occur. The Pocahontas group, or Lower Pottsville measures, has a thickness of 750 ft. in southern McDowell County, which decreases to about 350 ft. at Prince, on the New River, in Fayette County, 50 miles to the northeast. In this group, at various points, 12 coal beds have been found, of which 4 have been found in workable thickness and are now being mined. The New River group, or Middle Pottsville measures, is just above the Pocahontas group and varies in thickness from 1300 ft. in southern McDowell County to about 700 ft. along New River. At various places, 16 beds of smokeless coal have been identified in this group, of which 5 have been found in workable thickness and are now being mined. A general columnar section of these strata is shown on Fig. 2. The workable seams in these strata, arranged in geological sequence, are as shown on page 77.

Jan 1, 1932

By Michael A. Farrugia, Marcel Michaud

Remoteness of western Canadian coals has always been a barrier to the use of Canadian coals domestically. However, the world energy crunch, increased mining costs, stricter environmental controls, labor shortages, etc., are such that by 1980 a shortfall from traditional U.S. sources will prevail. Of necessity, Canada must then bring western coals east, requiring major new transportation initiatives. CN, CP Rail and Neptune Terminals are proceeding jointly with the design of a $30 million, fully automated, coal terminal at Thunder Bay. This element forms part of a $400 million integral transportation system to open new lines, improve trackage, acquire rolling stock and ships, for an initial capacity of 3 million tpy.

Jan 1, 1977

By A. J. Lanza

IN view of the immense amount of attention that silicosis has received in this country in the past few years, it is timely to review the status of the silicosis problem at present. Who gets silicosis? The individual exposed to the inhalation of free silica dust. The effects upon the individual depend upon the amount of silica inhaled or, in other words, the severity of exposure, rather than upon any special variation or susceptibility among those exposed. A considerable amount of study has been given to determine what may be the role of predisposing causes in the development of silicosis. To men-tion one instance, at the Picher Clinic careful records were kept of all previous illnesses of the men examined as well as abnormalities present at the time of examination. Examinations of over 27,000 miners were made during the years the clinic was conducted under governmental auspices-a very fair sized sample. The general conclusions from all this work were that the incidence of silicosis is not appreciably influenced by racial, structural, or other inherent qualities; it is influenced by acquired defects-particularly as these affect the lungs. That is to say, the average run of human beings respond similarly to similar exposures to silica dust and those whose lungs are already impaired by tuberculosis or other pulmonary disease acquire silicosis more readily than those with healthy lungs. It may well be, and there is some clinical evidence in support, that negroes acquire silicosis more readily than do whites, but the infection factor has not been so carefully studied among negroes exposed to silica as it has among whites. We know that tubercle infection is more prev-alent among negroes than among whites. There may be other factors concerned in the development of silicosis in negroes as compared to whites, but there is not available information upon which to base posi-tive statements.

Jan 1, 1937

By -1ng. B. Stampa

Introduction The author investigates the technology available for mine flow reversibility by reversing the operation of the main fans. Air bypassing, reversed sense of impeller rotation, and blade angle overturning are indentified as the most important means of achieving the flow reversibility. The axial flow fan is studied in some detail and the important variables of fan capacity, pressure, and efficiency are compared. Although some references to rotor velocity vectors are made (Fig. 1) the subject is not presented in detail in order to stay within the paper&apos;s intent. Background Where applicable, the requirements of mine air reversibility are an important part of the overall mining operation and should be given proper attention during mine ventilation design. The need for preventing mine fires in intake shafts is the reason that, for a long time, it has been a requirement in the Eastern European countries to have the feature of flow reversibility built into all main mine ventilators. In West German coal mines, this requirement does not apply since higher priority is given to safety precautions. Some metal and mineral mines in the United States are required to have ventilation currents readily reversible in order to fight underground fires as stipulated by the Title 30, Mineral Resources, and by the mining laws of some states. From an equipment performance standpoint, assuming the validity of other flow technical data, the question of the ventilator flow reversibility is of ten related to the actual scenario of the begining of an emergency. Most of the time, the system operates within half of the design capacity only and assuming that the equivalent mine orifice is stable only a quarter of the design pressure must be produced by the fan in any event. For such an operating point, the effective gradient and energy losses can be disregarded since fire emergencies usually last for only a short period of time. Equipment Comparison Fig. 2 depicts a ventilation arrangement wherein the flow can be reversed through changing the functional position of the dampers (slideplate or blade type). Both axial or radial fans can be applied in this case. For the mine exhaust operation, slide plates (guillotines) I and I11 are opened while plates I1 and IV are closed. The flow thus encounters two reroutings, which due to the reverse ventilation flow cannot be equipped with proper sheet metal deflectors, and therefore large pressure losses result during the air reroute operation. When the intake operation is considered, the dampers I and I11 are closed and the pair I1 and IV is opened. For this arrangement, almost the same flow conditions exist as for the foregoing exhaust mode; however, one additional corner must be applied. Even if a highly efficient ventilator is used in one of the previously described systems, the additional pressure losses resulting from air rerouting plus those encountered between the planes "0" and "00" significantly affect the overall energy requirements. Illustrated in Table 1 are the relative component losses of even flows for evaluation of various detour arrangements. From the aerodynamics standpoint, when these components are being designed, serious consideration must be given to the individual levels of the pressure loss coefficient, 5. Should there be an increase in the sum of the loss coefficients due to the particular component selection such that, for example, the net change is 1.2; the resulting pressure to be produced by the fan can be calculated from:

Jan 1, 1981