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By Games Slayter

NO reasonably well-informed person believes that the role of minerals, both metallic and nonmetallic, will be any less important in the future than it has been in the past. The contrary is true. Industrialization is based on mineral resources and man's ability to make use of them. A nation's standard of living, its political and military power, will continue to depend largely upon the extent to which it is industrialized-and minerals are the raw materials both of the machine and the power to run the machine. Minerals are also the raw materials of human health and stamina. The world's healthiest, most vigorous, and most long-lived peoples have always been those that have lived in regions of the globe where the earth is rich, and kept rich, in the phosphates and other minerals required for the production of the so-called growth foods. Only a small proportion of the land surface of the globe is suitable under any circumstances for large-scale production of these and other essential foods. As things stand, only a part of this land is capable of such production either because the soil minerals have been exhausted by cultivation, or because they never existed in the soil in sufficient quantities.

Jan 1, 1943

By John S. Marshall

The purpose of this paper is to present to the profession data and experience obtained over a period of 5 years in the operation of 87 shaker-conveyor units, and the production of 2,169,638 tons of run-of-mine anthracite coal. The data are not generally available and should be of interest to those concerned with the operation of anthracite and bituminous mines. To simplify comparison, a ton of 2000 lb. of raw run-of-mine coal has been used as the production unit throughout this paper. Mining Methods and Equipment The mine is in the vicinity of Wilkes-Barre, Pa., in the Buried Valley area of the Susquehanna River. Seven veins were mined, varying in thickness from 18 in. to 8 ft. The cover overlying the veins varied from 200 to 800 ft. Three-fourths of the 1200 acres of property was covered with from 100 to 200 ft. of water-bearing gravels and clays, which prevented first mining under rock cover less than 50 ft. and pillar removal under rock cover less than 100 ft. The property was opened by two shafts sunk on the mountainside to avoid the sand and gravel present in the river valley. The veins dipped an average of 5" from the shafts. In certain areas the veins were practically flat and in some places a reverse pitch as high as 5" was encountered for short distances. A daily output of 3800 tons was hoisted on a two-shifi; hoisting operation. The mining consisted of 75 per cent pillar removal, and of the total production in the later years of operation 80 per cent was loaded by shaker conveyors, chain conveyors and scraper loaders. Shaker conveyors accounted for 70 per cent of the total output; chain conveyors and scraper loaders for 10 per cent. All conveyor work was double-shifted and 25 per cent was triple-shifted. A total of 87 shaker-conveyor units were on the property, of which 75 were kept in operation at all times.

Jan 1, 1940

By John S. Marshall

The purpose of this paper is to present to the profession data and experience obtained over a period of 5 years in the operation of 87 shaker-conveyor units, and the production of 2,169,638 tons of run-of-mine anthracite coal. The data are not generally available and should be of interest to those concerned with the operation of anthracite and bituminous mines. To simplify comparison, a ton of 2000 lb. of raw run-of-mine coal has been used as the production unit throughout this paper. Mining Methods and Equipment The mine is in the vicinity of Wilkes-Barre, Pa., in the Buried Valley area of the Susquehanna River. Seven veins were mined, varying in thickness from 18 in. to 8 ft. The cover overlying the veins varied from 200 to 800 ft. Three-fourths of the 1200 acres of property was covered with from 100 to 200 ft. of water-bearing gravels and clays, which prevented first mining under rock cover less than 50 ft. and pillar removal under rock cover less than 100 ft. The property was opened by two shafts sunk on the mountainside to avoid the sand and gravel present in the river valley. The veins dipped an average of 5" from the shafts. In certain areas the veins were practically flat and in some places a reverse pitch as high as 5" was encountered for short distances. A daily output of 3800 tons was hoisted on a two-shifi; hoisting operation. The mining consisted of 75 per cent pillar removal, and of the total production in the later years of operation 80 per cent was loaded by shaker conveyors, chain conveyors and scraper loaders. Shaker conveyors accounted for 70 per cent of the total output; chain conveyors and scraper loaders for 10 per cent. All conveyor work was double-shifted and 25 per cent was triple-shifted. A total of 87 shaker-conveyor units were on the property, of which 75 were kept in operation at all times.

Jan 1, 1940

By John Birkinbine

The experience of the past five years has furnished opportunities to study and to partially explain the operation and some of the causes of the short blast of Warwick Furnace at Pottstown, Pennsylvania, as described in the paper read at the Lake Superior Meeting, in August, 1880.* The many inquiries concerning the details of furnace operation which followed the publication of the paper, and the expressed desire of furnace managers to be informed as to future operations of this plant, have encouraged the preparation of the following data. To make the record continuous, and add to the value of both contributions, the present paper commences where the review of a short blast of Warwick Furnace ended, and includes the data of another short blast of two months, and a subsequent continuous campaign of nearly five years, which, by comparison, we shall designate as "the long blast." Frequent visits to the plant in the five years, during which its operation could be studied, and correspondence with the manager, Mr. Edgar S. Cook, have given a knowledge as to general features and details, and the courtesy of the officers of the Company which has given free access to the books of record permits the presentation of data which it is believed will be welcomed by furnace managers. Interest will undoubtedly be increased by the remarkable results which Mr. Cook has achieved chiefly with anthracite coal in a furnace whose details of construction will scarcely be generally admitted as embodying advanced ideas, its height being but 55 feet, with a bosh diameter of 15 feet 6 inches, its bosh-slope comparatively flat, its crucible smaller than the average, and the heat of the blast never reaching 1000' Fahrenheit. At present writing the credit of the best anthracite blast-furnace practice recorded in our Transactions is conceded to be that pursued by Mr. Fackenthal, at the Durham furnace, and Mr. Cook, at War-

Jan 1, 1886

By C. M. Wayman, J. F. Breedis

The crystallography of the martensitic transformation in the Fe-30 Ni alloy was reinvestigated. The scatter in the habit Plane as determined from measurements on the mid-rib plane was smaller than reported previously, and the orientation relationship was determined from a martensite plate having known direction cosines. From agreement with crystallographic theory and from observations of the morphnlogy of the martensite plates, it is suggested that the mid-rib plane represents the starting region of the transformation. THE crystallography of the martensitic transformation in the iron-nickel system has been the subject of many investigations.&apos;-7 However, for iron alloys containing approximately 30 wt pct Ni, a wide scatter in the reported habit planes exists. In only one instance has a Laue determination of the orientation relationship been made, but the habit plane of the martensite plate studied was not determined experimentally. Because the most successful thermodynamic and kinetic analyses8 of the martensite transformation have been based on Fe-30 Ni alloys, it was considered desirable to reinvestigate the crystallographic features of the transformation in this alloy. Lieberman9 suggested that the most significant information for use in a theoretical treatment involves the determination of all the crystallographic observables from a single plate of the product phase. This was first done in essence by Greninger and Troiano10 for Fe-22 wt pct Ni-0.8 wt pct C, and is the method followed in this investigation. Some theoretical approaches11-13 in describing the martensitic transformation consider the overall transformation distortion to be comprised of a Bain distortion, or lattice deformation (PI, which produces the structure change, and a lattice invariant deformation (G), such as twinning, which together leave the interface or habit plane un-distorted. The incorporation of a rigid body rotation (R) results in a transformation distortion (E) such that the interface plane is left both undistorted and unrotated. In matrix notation, E = RPG where the lattice invariant deformation has "taken place" in the austenite. Bowles and Mackenzie14-16 in formulating a theoretical approach proposed an adjustable scaler quantity, denoted as 6, which represents a uniform dilation in the interface plane. This dilation acts to adjust the lattice spacings of the parent and product structures at the plane of contact. Using the dilation parameter as a variable, these authors were able to account for a continuous variation of habit planes in steels along a curve in the stereogram between {225}A and {259}A. she invariant line strain14 of the transformation is 6 RP, and the corresponding transformation distortion is E/6.) With the dilation parameter known, the invariant line is computed as the triple coincidence of the habit plane, the shear plane of the lattice invariant deformation, and the initial Bain cone.14 The contraction axis for the Bain distortion can be taken9&apos;17 as [010]A, and the correct permutation of the habit plane indices used. Kelly and Nutting 18 have established the shear plane (of twinning) in the tetragonal martensite of Fe-22 Ni-0.8C to be {112)M by means of transmission electron microscopy. The reasonable shear system is then {112}M which is equivalent to {011}A <01l>A. The as-

Jan 1, 1962

[A Aall, Christian H., M49 D Mgr, American Metal Climax, Inc., 61 Broadway, New York 6, N.Y. Aaroe, S. L., A54 B Sales Mgr, The W. S. Tyler Co. of Can. Ltd., St. Catharines, Ont., Canada Abalos, Teodulo Q., J57 A Foreman, Lepanto Cons. Ming. Co., Mt. Province, Philippines Abbass, G., M55 P. O. Box 680, Hong Kong, China . Abbitt, James M., A54 770 Sole Blvd., Bartow, Fla. Abdullah, M. Y., J53 Dept. of Mines Res. Div., P. O. Box 1016, Tiger Lane, Ipoh Perak, Malaya Able, John F., Jr., J53 A Ming Engr, USA SIPRE 1215 Washington Ave., Wilmette, Ill. Abell, Leo M., M31 B Conct Supt, Phelps Dodge Corp., Morenci, Ariz. Abenius, Hakan Wilhelm, M52 A Mng Dir, Stora Kopparbergs Bergslags AB, Falun, Sweden Abenoroth, Harry F., A57 Natl. Carbon India Ltd., 18A Brabourne Rd., Box 217, Calcutta, India Abercrombie, Harry W., J56 1645 Camphor Dr., Lakeland, Fla.]

Jan 1, 1961

By Oliver Bowles

PRICES of quartz sold in the United States in 1938 ranged from $1.15 to $36,000 a ton. This startling variation was due simply to the differences between glass sand and rock - crystal, materials that are virtually the same in composition. It illustrates the complexities in qualities, uses, and values of industrial minerals. It is generally stated that mineral deposits are exhaustible resources which can never be replaced when depleted. This statement is no longer true with respect to certain industrial minerals now recovered in increasing quantities from that most inexhaustible of all sources of supply, the ocean. Salt has long been obtained from sea water, particularly on the Pacific Coast, and reference was made a year ago to the large bromine recovery plant at Wilmington, N. C. Now the Westvaco Chlorine Products Corp. is operating a large plant at Newark, Calif., having a capacity sufficient to recover daily from the ocean raw materials that will produce 60 tons of various magnesium products. Lime is used to precipitate the magnesia from salt-water bittern. The products are chiefly magnesium carbonate for chemical uses, periclase and other refractories, and oxychloride cement for flooring.

Jan 1, 1939

By BLABIEY STEVENS

THE object of this paper is to present a theory of the formation of fissures which seems to be supported by all available data. The investigation is, in the main, an exact one, and irregularities of the rock-structure are generalized. From the conclusions drawn a general classification of fissures is made, and this should be of practical use in the comparison of mineral deposits. The theory also throws some light on the equilibrium of the earth&apos;s crust and the depth of action of surface-waters. Incidentally, additions are made to the theory of earthquake-slips. Former Theory. It has been previously pointed out that the &apos;"jointing" of homogeneous rocks occurs in two sets of planes, at right angles to one another, and making angles of 45° with the directions of greatest and least principal stress.&apos; The theory by which this fact is explained does not, however, account for fissures and faults, because normal faults are usually more vertically, and reverse faults more horizontally, inclined than 45°. The "jointing law" is based on the following axiom: "In a homogeneous mass under pressure, rupture must take place on the lines of maximum tangential stress." 2 This law applies, with fair approximation, to the tests on jointing which can be made in the laboratory, and as nearly as can be expected to the natural jointing-planes in rock-masses, where the stresses are not very great; but it entirely fails to account for fissure-fractures.

Aug 1, 1909

By L. M. Howe, J. Gordon Parr, E. Saarema

THE decreasing solid solubility limit at the titanium-rich end of the Ti-Cu constitutional diagram,&apos; Fig. 1, suggests the possibility that titanium-rich alloys may be age-hardenable. However, results obtained by Holden et al. reproduced in Fig. 2 show that after quenching from 790°C the age-hardening of an alloy containing 1.7 pet Cu is very slight, while a 0.8 pet alloy decreases in hardness, during heat treatment at 400°C. It was believed possible that powder samples of alloys might show different results from the lump samples used by Holden et al., since small particles are often more sensitive to diffusion processes than larger samples. Consequently, a 1.9 pet (by weight) Cu alloy was made by the technique of levitation melting," checked for homogeneity, and filings of 48-65 Tyler screen size were cut from it for aging experiments. The filings were helium-quenched (by a technique described previously)&apos; from 790°C, sealed in vacuo in silica tubes and aged at temperatures of 400°, 450°, and 500°C. Microscopic and X-ray examinations failed to reveal any structural changes during the aging process. This is not too surprising since early stages of precipitation do not usually manifest themselves in ways that are readily detectable by metallographic methods, and the extent of precipitation on overaging a 1.9 pet alloy is very small. However, hardness readings (taken on a Bergs-man micro hardness tester, using a 25 g load applied to a diamond indenter) do show a hardness peak at each aging temperature, Fig. 3. About a dozen readings were taken on each heat treated sample; the highest reading and the lowest reading were ignored and an average taken of the remaining values. In

Jan 1, 1957

By AIME AIME

Russell B. Paul, Chairman of the Special Committee on Divisional Relationships, presented the interim report of his Committee which was published in the September, 1945, issue of MINING AND METALLURGY and a copy of which is incorporated in the official minutes of this meeting. The following specific actions were taken by the Board of Directors in relation to the six recommendations in the Summary of the Report as published: (1) Mr. Haider, Chairman of the Petroleum Division, presented the formal request of his Executive Committee that the headquarters of the new Secretary of the Petroleum Division be established in Dallas, Texas. After lengthy discussion in which the point was made that the success of this program would depend upon effective co-operation between the Chairman of the Division, the Secretary of the Petroleum Division, and the Secretary of the Institute, the request was approved. The new secretary will be selected by the Division and will be employed by the Institute in the capacity of an Assistant Secretary. (2) The following item in the report was approved:

Jan 1, 1945

By Charles E. Lundin, Albert S. Yamamoto

The phase eqrtilibria of the binavy system, Nb-Al, weye established by metallographic, X-ray diffraction, hardness, and thermal-analysis techniques. Three intermediate phases were identijied, Nb3.41, Nb2Al, and NbAl,. The N6,Al and NbAl phases have extensiue ranges of solubility. The maximum solubility of aluminum in niobiuitz is 23.0 i0.5 at. pct (8.0 wt pct) A1 clt 1960" i 10°C. The maximumt solinbility oj niobiunz in alunzinutz is less than 0.3 i0.1 at. pct (1.0 tut pct) fib. The A&apos;h,Al and Nbgl interinediate phases forr peritectically at 1960° 1UcC and 1870" i 1U9C, respectively, zuhereas the NbAl, phase fornts congvuently fio11t the melt at 1605" 10°C. A eutectic reaction oc- The intent of an over-all research program was to develop niobium-aluminide-base coatings for refractory metals since these coatings exhibit useful high- curs at 57.8 i 0.5 at. pct (28.5 wt pet) A1 and 1500" 10°C between NbAl and NbAl3. The aluminum-rich region of the system has a peritectic reaction where liquid reacts with NbAL, to form aluminum solid solution. Structural and lattice -pavarneter data /or the intermediate phases were determined by X-ray diffraction analyses. An unusual transfov?nation having charactevistics of nucleation-growth and martensitic nleclanisils occurs on cooling the niobium solid solution in the region from 18 at. pct (6 wt pct) Al to 23 at. pct (8 wt pct) Al. The transfornzed product on rapid quenchirzg is an acicular NhAl. temperature resistance to oxidation. As a prerequisite, the phase equilibria of the system, Nb-A1, were necessary. A literature survey revealed that the phase-equilibria data reported by various investigators were rather incomplete and conflicting. Therefore a complete determination of the system was undertaken by the authors from 0 to 100 pct of solute addition.&apos; In 1959 Savitskiyz presented a tentative diagram which

Jan 1, 1967

By W. E. Rudolph, R. E. Baylor

DUE to its location in the Atacama Desert, one of the most barren of the earth&apos;s surfaces, Chuquicamata&apos;s water supply presents unusual problems. Yearly rain-fall averages less than one tenth of an inch at the plant. However, there are summer showers above 12,000 ft in the Cordillera to the east, the resulting run-off flowing through old river valleys buried beneath more recent volcanic formations, to be impounded within sediment-filled basins. This water emerges at springs where the outlets of these basins are blocked by lava flows, and here are formed the small streams which feed the only important river of the region, the Rio Loa. Chuquicamata&apos;s water is obtained from these springs and rivulets. [ ] The map above indicates four pipe lines from which potable and industrial water are supplied. Potable water, amounting to 4500 metric tons per day, is conveyed in the Toconce pipe line from springs 59 miles due east of Chuquicamata. This water is used not only for drinking, but also for boilers and other needs requiring high quality. For industrial water at the oxide plant, there are two 12-in. pipe lines from the Rio San Pedro, carrying a total of 17,000 metric tons per day of slightly brackish water. This water is at present used mainly for leaching and for hygienic purposes. Water Source Found For the present and future needs of the sulphide plant, it was calculated that at least 32,000 metric tons per day of make-up water would be required. For this purpose, a pipe line of 44 miles length was constructed to bring in the entire flow of the Arroyo Salado, one of the eastern tributaries of the Loa. The salt content of this water is so high (over 5000 parts per million of solubles, mostly chlorides) that it is highly detrimental to farming, and the Chilean Government had been studying projects to separate these waters from others of the Loa system in order to improve agricultural conditions in the fertile valley of Calama. So it happened that the Government was willing to award rights to the Arroyo Salado waters under agreement whereby the Mining Company removes waters from the Rio Loa system above Calama for all time. The outlet of these waters, after serving their purpose at the new concentrator and leaving the plant in tailing, is the Salar de Talabre, an old salt lake which presents fully ten square miles of surface to serve as an evaporating pan, the outlets having now been blocked by dams. Here the dry climate of Chuquicamata is a favourable factor, evaporation averaging slightly above 1/4 in. per day. The Toconce and San Pedro pipe lines have been functioning from 26 to 34 years, and through the use of special cleaning tools which were developed at the plant, as well as deaeration of the more active Toconce water, these pipes are now maintained at capacities which do not diminish as years go on. Constructing the Dam The Arroyo Salado pipe line design and construction involved certain special and interesting features, and inasmuch as this line and its intake works are solely for the needs of the new sulphide plant, more detailed description is given. The waters are impounded at a gravity dam constructed of concrete to a height of 100 ft above the river bed, keyed into the precipitous Dacite walls of the narrow canyon (barely 6-ft wide at the bottom, only 25-ft width at 50 ft above). A small secondary dam was built 100 ft down stream from the main dam, providing a pool of 15-ft depth to protect the main structure from flood flows over the spill-way during the rainy season. A system of four 36-in. syphons was designed for discharging these flood waters from the lower depths of the lake, in order to avoid eventual sedimentation behind the dam. The lake has a length of 3300 ft, and its water level is controlled by an adjustable spillway permitting draw-down of eighty inches, amounting to 41,000 metric tons of available capacity. This regulation is necessary because of wide fluctuations in stream flow between day and night due to freezing of feeders. During the construction of the dam the entire river flow was handled within a 36-in. pipe line some 2000 ft in length. As the excavations proceeded

Jan 1, 1952

Eckley B. Coxe, Drifton, Pa.: Anthracite coal differs from other fuels in its greater solidity. It does not burn like coke or wood, or like bituminous and coking coals, which become more or less spongy when heated, so that combustion can take place to some

Jan 1, 1892

By Carle Hayward

SULPHUR has long been one of the banes of the steel manufacturer and often no effort and expense have been spared in order to reduce it to a small per cent. in the finished product. This condition is clue to a general conviction that in many cases where steels have failed in service, sulphur has been the cause. But there has been a growing feeling in recent years that the verdict against sulphur has been unnecessarily severe. In cases of segregation it was present in augmented amounts along with other impurities, but it had not caused the segregation. High sulphur in pig iron is caused by poor furnace conditions and the sulphur is merely one indication of an iron that has not been properly reduced. No amount of subsequent treatment under oxidizing conditions in the open-hearth furnace can remedy the defects, although the per cent. of sulphur may be considerably reduced. In other words, the causes of bad steel can frequently be traced back to bad pig iron, and sulphur is merely one indication that the pig iron is bad. The writer recently visited a steel plant where a mass of evidence had been accumulated which substantiated this fact, and the superintendent was .emphatic in stating that high sulphur was not harmful provided the steel was not otherwise poor due to insufficient reduction in the blast furnace. The presence of a moderate amount of sulphur is desirable from the standpoint of the man who machines the steel. The low-sulphur material drags and the production of a smooth surface is very difficult. A slight increase in sulphur enables the machinist to produce a smooth surface without difficulty. Since, therefore, such large quantities of steel are subjected to machining, it becomes highly important that the sulphur controversy should be settled, and if its presence is proved to be harmless the ban on it should be lifted.

Jan 10, 1916

By J. A. Morgan, R. E. Lund, D. E. Warnes

The design, operation, and performance of an integrated pilot plant for recovering zinc and copper from a complex sulfide ore are described. Metallurqical processing comprised selective sulfate roasting in a fluid bed, leaching in a weak sulfuric acid solution, recovering copper by cementation with scrap iron, air oxidation to remove soluble iron, more-or-less conventional Purification for removal of impurities deleterious to zinc electrolysis, recovering zinc by electrolyzing solutions, and recycling spent electrolyte to the sulfating roaster. The effects of several feed preparation and roasting variables on the sulfation response of zinc, copper, and iron are set forth, together with the electrolyzing characteristics of the zinc solutions. SEVERAL years ago, St. Joseph Lead Co. undertook a research program to develop procedures for winning values from a complex ore. The deposit was a massive sulfide ore body containing upwards of 65 pct pyrite. Gangue, composed principally of quartz, calcite, dolomite, and alumina, comprised only 10 to 20 pct of the mineralized zones. Zinc, lead, and copper values were variable, but averaged about 6.7 pct Zn, 2.6 pct Pb, and 0.4 pct Cu. Early in the exploratory test work, it was established that the intimate physical association of the minerals in the ore made beneficiation by conventional flotation procedures difficult. Accordingly, in addition to a comprehensive program to investigate physical methods of ore beneficiation, a research program was established to develop a direct metallurgical method for processing the ore. After surveying a number of potential methods for recovery of both metal and sulfur values, effort was concentrated on the development of a sulfate roasting process. It is the pilot-plant development of the sulfate roasting-solution purification—electrolyzing procedures which forms the subject of the present paper. PREPILOT PLANT INVESTIGATIONS From a thermodynamic viewpoint, it is feasible to sulfate selectively a host of metals, including Zn, Cu, Pb, Cd, Ni, and Co, in an iron sulfide ore without sulfating iron. Of course, a number of conditions such as roasting temperature and gas composition must be controlled. In addition, as first disclosed by Hybinette,&apos; an alkali metal salt is frequently beneficial in promoting the sulfation reactions. The application of fluid bed roasting to the sulfate roasting of nonferrous metals was pioneered by Stephens of Battelle Memorial Institute.&apos; Subsequently, Falconbridge Nickel Mines, Ltd. developed an unique process for selectively sulfating and recovering nickel, copper, and cobalt from their nickeliferous pyrrhotite, which recently has been described in an excellent paper by Thornhill.3 One of the unique features of the Falconbridge process is that the feed is introduced into the fluidized roasting furnace in a manner such that the feed slurry is broken up into droplets and the droplets lose their moisture before contacting the surface of the fluidized bed. The dried droplets do not tend to cake or fuse, but rather remain as discrete agglomerates of concentrate particles and roast without significant degradation. St. Joseph, aware that Battelle Memorial Institute had participated in the development of the Falconbridge process, obtained permission from Falcon-

Jan 1, 1962

By B. M. Larsen, T. E. Brower

A perspective is presented of the steel plant sulphur distribution and elimination problem from coal to liquid steel ready for teeming, giving distributions of sulphur over a range of coke sulphur content, and some methods of sulphur control, in the blast furnace, external desulphuriza-tion between blast furnace and open hearth, distribution between fuel, slag, and metal, and methods and limitations of control of sulphur in the open hearth furnace. AS a part of the 1951 AIME symposium on sulphur in steelmaking, it was thought that a discussion of the distribution of sulphur throughout the whole series of operations, from coal and ore to finished steel ingots, might have some value in giving a perspective on the whole problem. The following discussion is an attempt to present such an overall picture. The order is that of the actual plant operations, beginning with a very brief consideration of the coking process. Sulphur in Coal and Coke Since by far the largest source of sulphur entering the steelmaking cycle is in the coal used to make coke for the blast furnace, it would seem reasonable to eliminate some of it, either from the coal, or the coke, or during the coking process. This has appeared impracticable up to the present, at least, for two main reasons: the low activity of the organic sulphur in either coal or coke, and because of price limitations involved in treating a low cost material such as coke. A variable portion, usually M or less, of the sulphur is present in coal in the form of pyrites or similar compounds, and a large part of this sulphur may be removed in the coal washery. Most of the sulphur, however, is normally present as "organic" sulphur, intimately associated with the coal structure. Its distribution prevents any separation by mechanical means. Its low activity makes improbable &apos; any rapid chemical removal, although hydrogen will remove sulphur from both coal and coke. Thus, prolonged recirculation of coke oven gas in the coking process would tend to leave a smaller percentage of the total sulphur in the coke residue. Table I shows a typical distribution of sulphur from coal into products in the coking process. As the sulphur in the coal increases, the sulphur in the coke tends to increase in about the same proportion. Sulphur in the Blast Furnace The best picture of the situation in the blast furnace is provided by a sulphur balance of raw materials entering, and of products leaving, the furnace. The difficulties in accurate weighing and sampling of the variable solid materials entering this process, and the number of hours required for the raw materials to descend through the furnace under variable operating conditions, make it difficult to obtain an accurate balance. However, balances made over periods of weeks or months tend to average out some of these uncertainties. Table I1 presents three typical sulphur balances similar to a number that the writers have calculated. In most of these the slag volume calculated from the sulphur balance is, in some instances more, and in other instances less, than the value corresponding to the best input and output balances of the other slag constituents (lime, silica, alumina, etc.). Probably the greatest source of error in these calculations is the sulphur content of the slag. Despite some possible inaccuracies the balances of Table II show rather definitely the following points: 1—That 87 to 95 pct of the total sulphur input is in the coke and 95 to 97 pct of the total sulphur output is in the slag. Also, that if any sulphur leaves the furnace with the gas it is relatively small, amounting to a possible 1 pct or less. 2—At the lower sulphur coke level of 0.86 pct the total amount of sulphur charged is 15 Ib of sulphur per ton iron increasing to 26 lb per ton at the higher sulphur, intimately associated with the coal struc-rare burdens containing sulphur-rich ores will the total sulphur burden fail to be nearly proportional to the content in the coke used. 3—The 7 to 9 pct of the total sulphur input from the limestone of furnaces B and C is due to the relatively high sulphur content of the stone, 0.226 and 0.265 pct, respectively. In the case of furnace A, the sulphur content of the limestone was only 0.06 pct which resulted in only 3 pct of the total sulphur input coming from this source. It is rather interesting to compare the sulphur balances of a typical ferromanganese furnace with

Jan 1, 1952

By C. W. Purington, H. B. C. Nitze

The Kotchkar mining-district, known as the Kotchkar System, is situated in the Orenburg government, in Eastern Russia, in the great plateau- or steppe-country immediately adjacent to the eastern slopes of the Ural mountains. Its longer dimension is north and south about 27 miles, while in the widest (northern) portion the distance across is 22 miles. The district lies 40 miles southwest of Miass, a station of the Trans-Siberian railway, and half-way between this village and that of Troitsk. It contains also the not inconsiderable settlements of Kotchkar and Eosobrodskoit., the last a Kossak town. So far as explored, the district is known to contain from 300 to 400 auriferous veins. ? Although the topography is generally of an exceedingly flat character, the country is traversed by several small rivulets running from west to east, and flowing into the Sanarka and Kamenka rivers. Since 1844 placers have been worked in the district, but it was not until 1867 that veins were discovered, and these have been worked only since the impoverishment of the alluvial deposits. The deposit with which this paper especially deals is known as the Ouspensky mine, one of twelve owned by MM. de Zelenkoff & Cie., five of which are situated in the Kotchkar System. An opportunity of visiting this mine was afforded the writers, in eonnection with an excursion of the VII. session of the International Geological Congress; held in Russia during the summer of 1897. The mine is reached by a 40-mile &apos;(troika" journey from the railway station of Bichkil, itself

Jan 1, 1899

By George S. Rice

Instantaneous outbursts of gas in underground workings are similar in effect to great blasts of explosives, but without heat effects. Fortunately they occur only in a few localities in exceptional coal fields of the world and under exceptional geologic conditions. Such violent outbursts should not be compared with the ordinary outflows of gas into coal mines or with explosions of gas and dust. The phenomenon is essentially one of gas held in circumscribed areas under very high pressure and as a mining face or heading approaches such a place and the mining excavation sufficiently weakens the natural surrounding dam or shell, this is burst by the pressure just as by the discharge of a mass of explosives. The outbursts, so far as known to the writer, have been confined to coal beds and have not occurred in other mineral deposits. Hence, as discussed later, he believes that it is the relation of the gases to the coal material that is a factor in conjunction with others in producing conditions which permit the phenomena to occur when mining is done. The gases in most instantaneous outbursts have been hydrocarbon gases but in two coal fields they have been definitely found to be chiefly or wholly carbon dioxide. One of these localities in which CO2 outbursts have been frequent and have caused loss of life is in the southern part of the Gard basin, near Alais, in southern France. In 1919, the writer visited the Fontaine mine in this field the day following the occurrence of an outburst by which two men were smothered by carbon dioxide and fine dust. After 24 hr. the CO2 had not been entirely removed by the ventilating current. Occasional outbursts of CO2 have also occurred in other small basins in France, causing accidents, as in Puy-de-Dome. However, mines in the Gard basin, north of those in which CO2 outbursts have occurred, have experienced hydrocarbon gas outbursts. In the Lower Silesian coal field, Germany, there is a group of mines which have been subject to many instantaneous CO2 outbursts, some causing serious accidents, but the greatest disaster occurred recently

Jan 1, 1931

By W. C. PAGE

ALL mine ore here must be concentrated before shipment, which involves selective flotation. Three products are made: lead, zinc, and pyrite concentrates. The equipment and practice are so well outlined in the accompanying flow sheet that only a few amplifying remarks are necessary. After crushing at the mine to approximately 4 in., the ore is carried by aerial tram about 4 1/2, mi. to the mill. From 1800 to 2000 metric tons per day, excluding Sundays, is thus transported. At the mill, the buckets are hand-tipped into the bins. After crushing to pass a ¼ -in. square opening, the ore may be stored in reserve crushed-ore bins or ball-mill feed bins as desired. Total bin capacity at the mill is about 4000 metric tons. The concentrator is arranged in three independent grinding sections, each fed from its own storage bin. Grinding operations follow standard practice, with ball mills in single-stage closed circuit with classifiers, the overflow from which is delivered to the lead circuit rougher flotation machines,

Jan 1, 1936

By John M. Campbell, W. E. Portman

A series of correlating charts have been prepared to enable the field engineer to predict the amount of stock tank fluid produced by stabilization of first stage separator fluid. The charts shown are particularly convenient for estimating the relative effect of temperature and pressure in economic studies. Further work is planned to investigate the factors affecting stage separation&apos; yo supplement this preliminary work. The direct correlation of separation variables is dictated as a convenience to eliminate the trial and error calculation necessary when using the standard flash vaporization equations. These are obtained by a material balance using the equilibrium vaporization ratio K, and have appeared in the following, or similar. form many times in the literature. where La = mols of any component n in the separator liquid Ko — equilibrium vaporization ratio for any component at separator pressure and temperature Fa = mols of any component n in the feed Vn = mols of any component n in the vapor V = total mols of separator vapor These equations may of course be effectively and accurately solved by use of digital and analog computers. However, these are not usually readily available in field offices and the need often arises for fast results during routine operational problems. Consequently, many methods have been devised in an attempt to simplify the basic calculation.2-14 The methods proposed are generally satisfactory, but do not eliminate trial and error methods and the necessity of subsequent calculations to obtain the answer in volume terms. The purpose of this work. therefore, was to develop accurate charts that would give the result directly in gallons. When using stage separation the recoveries would be lower than those shown, for stabilization usually gives about 95 + per cent recovery on the potential stock tank components in the separator liquid. The recoveries are shown in terms of a 14 psi Reid Vapor Pressure product which has a true vapor pressure of 14.7 psia at 98°F. Consequently, this product represents a stable stock tank liquid under normal storage conditions. To estimate natural gasoline re- coveries by stabilization the volume shown may be multiplied by 1.20 for 18 Ib RVP natural gasoline and 1.36 for 26 lb RVP natural gasoline. PROCEDURE The five analyses shown in Table 1 were chosen since they were representative of wellstream compositions often found and had almost the same propanes plus liquid content, expressed in gallons per 1,000 std. cu ft of wellstream. The distribution of the components though is widely variant. A temperature range of 0-100°F. and a pressure range of 200 to 1,250 psia was chosen inasmuch as most separations fall within these limits The K values used were modified values of those originally published by Katz and Hachmuth.&apos; In the range chosen these have generally given good results when predicting fluid recoveries. The choice was somewhat arbitrary, for the authors&apos; personal experience has shown that

Jan 1, 1957