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By Alden B. Greninger

Many recent contributions in the field of theoretical metallography have been concerned with crystallographic definitions or descriptions of various phenomena. The lattice orientation of the crystal being studied must be known before such definitions can be made. Practically all the metallic crystals used do not have crystal faces—other than the unknown "plane of polish"; hence resort must be had to other means of determining orientations. Some use has been made of etch-figures, slip bands, optical properties, etc., for this purpose; but for accurate results, and because of the fewer restrictions to their use, the various X-ray methods have been the most satisfactory.' The methods that have been used may be divided into two general classes: (1) oscillation methods, and (2) transmission Laue methods.² The disadvantages of these methods are numerous, chief among them being the time required to determine a complete orientation. No method proposed so far is wholly satisfactory for determining the orientation of a small macrograin within a polycrystalline aggregate.

Jan 1, 1935

By Clarence Wright

THE average lead and zinc content of the ores mined and milled in the Joplin district is low as compared with that of other lead and zinc deposits throughout the United States. Because of this fact and of the high grade of the concentrates, the ratio of concentration is unusually high. Most of the operating companies have aimed to produce as high a grade of concentrate as possible in order to obtain the highest price for their product. They have done this at the sacrifice of the percentage of zinc recovered. The average percentage of zinc recovery is from 60 to 70 per cent. For percentage of recoveries from actual mill tests, see tables. The concentration is commonly effected by crushing the ore to h-in. (12.7 mm.) size and subjecting it to a roughing and cleaning over two or more large Cooley jigs, of the Harz type, the finer sands being treated on the usual types of sand and slime tables. The treatment of slimes by flotation is being tested at several of the larger mills. COARSE CONCENTRATION When the ore. has been hoisted from the mine it is dumped onto a chute leading to a grizzly made of iron bars, usually heavy rails, spaced 4 to -5 in. (101 to 127 mm.) apart. The oversize is caught on the grizzly, the mineralized pieces are broken with sledge hammers so as to pass through, and the boulders that are practically barren are sorted out by hand and trammed to the waste-rock pile. The proportion of waste rock thus sorted out varies from a fraction of 1 per cent to 15 per cent., depending on the nature of the rock.

Jan 10, 1917

By Oscar Lee

RESERVE Mining Co., owned jointly, by Armco Steel Corp. and Republic Steel Corp., controls 9% miles of magnetic taconite ore on the eastern end of the Mesabi Range lying between Mesaba and Birch Lake. The ore reserve which can be mined by open pit methods is conservatively estimated at 1.5 billion tons under a shallow, loose overburden requiring no rock stripping. Taconite now being mined is in the upper cherty horizon. It has been metamorphosed by the intrusion of gabbro and most of the iron oxide is altered to magnetite. The nonmagnetic portion of the iron is practically all in the form of silicate. The magnetic iron is finely disseminated through the rock and requires grinding -150 mesh for clean liberation of the iron oxide from the gangue.

Jan 5, 1954

By R. T. DeHoff

Saltykov's meihod for the determination of the size distribution of spherical particles from measurements made on random plane sections is extended to particle shapes which are not equiaxed. The analysis permits the deterrrzination of the size distribution of aggregates of ellipsoidol particles provided all particles have tlze same shape. V ARIOUS solutions to the problem of the determination of the size distribution of spherical particles from measurements made on random plane sections have appeared in the literature since the original solution was first published by Scheill in 1934. Solutions have been offered utilizing measurements of the diameter1y2 or area3'4 of circular sections that result when a test plane is passed through an aggregate of spherical particles, and the length of intercept5y6 that results when a test line is passed through the aggregate. To the best of the knowledge of the author, no method has been proposed which permits the determination of the size distribution of particles that are not equiaxed. The purpose of this paper is to develop a method which permits the determination of the size distribution for a specific class of particle shapes which cover awide range of axial ratios; namely, ellipsoids of revolution. In the interests of simplicity, it is necessary to limit consideration to aggregates of particles which are all similar in shape; that is, the ellipses which, by rotation about one of their axes, generate the particles, all have the same axial ratio. The size distribution, which is to be determined by the analysis, is arbitrary. It is also assumed that the section, or sections, upon which the measurements are made represent the average of all possible sections through the structure. This ideal condition can be approached as nearly as may be desired by observing a sufficient area of sections taken at random positions in a randomly oriented aggregate, and in random positions and orientations in an aggregate that has one or more axes of orientation. SIZE DISTRIBUTION FOR PROLATE ELLIPSOIDS Consider an aggregate of cigar shaped, geometrically similar particles (prolate ellipsoids), dispersed in an opaque matrix. Let the size of each particle be specified by the length of its minor axis, By Fig. 1, and the shape, q, be defined by the ratio of its minor to its major axis, B/A. Let the index, j, refer to dimensions of ellipsoids in a particular size class, and the index, i, to dimensions of an ellipse resulting from the intersection of a test plane and a particle. The size of such a section is specified by the length of its minor axis, b. Let the particles be arbitrarily divided into k size classes, and let the increment of size between each class be A, where A = &,$, B,,,being the minor axis of the ellipsoids in the largest size class. Since all sections are represented on the test plane, the minor axis of the largest of the sections observed, results from the intersection of the test plane with an ellipsoid in the largest size class at its center, and is therefore equal to Bmax Thus the increment, A = The final result will consist of the k numbers which represent the number of particles in each size class per unit volume of aggregate:

Jan 1, 1962

By L. D. Fetterolf, G. T. Mahler, W. M. Peirce, R. K. Waring

UNTIL recently, the only commercial method of producing magnesium has been fused salt electrolysis, despite a considerable amount of experimental work on the direct reduction of magnesium oxide. In this early work, a number of investigators demonstrated that magnesium oxide would react with any one of several reducing agents to produce magnesium vapor, but that the mixture of magnesium oxide and reducing agent must be heated to a high temperature even in a vacuum. In particular, it was demonstrated that silicon and aluminum are effective reducing agents and that when either of these is used the magnesium vapor can be condensed to massive metal. In spite of this knowledge, no successful commercial reduction process was developed, at least in this country, because of the operational difficulties imposed by the requirement of high temperature and vacuum and because of the high cost of the reducing agent. The outstanding experimental work done by Pidgeon, who carefully reviewed all of the known methods of making magnesium and then selected ferrosilicon reduction, overcame some of the operational difficulties. The process developed by Pidgeon comprises briquetting a mixture of calcined dolomite and reducing agent, pre-heating the briquets to about 600°C. in an externally heated alloy retort, reducing the MgO with silicon in an evacuated horizontal alloy retort heated externally to about 1150°C., and condensing the magnesium vapor to a solid in a cold extension of the retort projecting beyond the fr6nt of the furnace. A major advantage of Pidgeon's process is that the retort is not cooled down between charges. Earlier investigators had used apparatus that necessitated cooling and reheating of the retort between charges. Early in January 1942, when the War Production Board was considering the Pidgeon process for quickly increasing the production of magnesium, The New Jersey Zinc Co. offered to have a research investigation of the process carried on, with the hope of assisting in its further development. This offer was accepted, and a two-retort pilot plant was built at Palmerton and put in operation on Jan. 29, 1942. An intensive investigation of certain phases of the process was carried out in the ensuing four months. This investigation was completed in the last part of May 1942, and the pilot plant was then shut down. Subsequently, The New Jersey Zinc Co. was requested to have further development work done on the process, and a more extensive investigation was undertaken in August I942 under contract with the War production Board and under the super- vision of the War Metallurgy Committee. This paper describes the phases of the investigation that are thought to .be

Jan 1, 1944

By Roshan B. Bhappu, Ronald J. Roman, Dexter H. Reynolds

The reaction between manganese dioxide and molybdenite in a water- sulfuric acid medium was studied at atmospheric pressure and from 25° to 103°C. Both solids are dissolved to give, as final products in solution, Mo(VI), S(VI), and Mn(II). Stoichiornetric relationshifis are most simply represented by the equation MoSz + YMnO2 + 15H+ - HMoO-4 + 2HSO-4 + 9Mn++ + 6H20 The rate of this reaction was found to be dependent on temperature, manganese sulfate concentration, acid concentration, and, in a complex way, the surface areas of both manganese dioxide and molybdenite. Both the total amount of manganese dioxide and tnolybdenite and the ratio of manganese dioxide to molybdenite played a part in determining the over-all rate of the reaction. The effect of acid concentration on the rate was also dependent on the ratio of manganese dioxide to molybdenite. SEVERAL reports in the literature describe reactions using manganese dioxide in water-sulfuric acid media as an oxidizing agent. Nearly all previous work describes interactions between an ion and manganese dioxide, such as the oxidation of ferrous to ferric iron,' or the interaction between a molecule in solution and manganese dioxide, such as the oxidation of toluene to benzaldehyde.2 Only a few references describe the use of solid manganese dioxide to oxidize another solid in sulfuric acid. One report3 states, "Manganese dioxide in dilute sulfuric acid solutions had no beneficial effect (in leaching copper from chalcopyrite), but in more concentrated acid solutions the rate of solution was about the same as that for permanganate." In 1902 a German patent4 described a process which used manganese dioxide in strong (50°Be') sulfuric acid for oxidizing chromite ores to soluble iron and chromium salts. The process consisted of mixing the sulfuric acid with the finely ground ore and manganese dioxide. The pulp was then heated to about 120°C and an electric current passed through it. The manganese dioxide apparently dissolved in the strong, hot acid solution, oxidized the iron and chromium and itself was reduced to manganese 11! then was oxidized at the anode back to manganese dioxide, and the cycle repeated. An extensive study5 of the kinetics of the oxidation of pyrite with manganese dioxide indicated that in dilute sulfuric acid solutions (about 0.1 N) and at elevated pressures the dissolution of the pyrite was brought about by air oxidation of the iron to ferrous sulfate, followed by manganese dioxide oxidation of the ferrous sulfate to ferric sulfate. The mechanism proposed by the authors involved the oxidation of the iron at the surface of the manganese dioxide. Also! the oxidation of sulfide sulfur to sulfate was credited to the oxygen in the air, not to the manganese dioxide. Since the metallurgical literature is lacking in information concerning the use of one solid to oxidize another solid. the current investigation was under-

Jan 1, 1965

By J. N. Hobstetter, J. J. Becker

Slip lines in deformed copper single crystals exhibit very pronounced cross-slip and clustering if the stress axes are not too far from <100>. This orientation dependence seems to persist whether or not attempts are made to reduce the bending constraints of ordinary mechanical tests. THE purpose of this paper is to report some microscopic features of the plastic deformation of copper single crystals. The material used was OFHC copper obtained from the American Brass Co. Single crystals about 3 in. long and 5/16 in. in diameter were made by lowering graphite crucibles containing this copper out of the hot zone of a vertical tubular furnace at a rate of 1 1/4 in. per hr. The furnace atmosphere was nitrogen. The orientations of the crystals were determined by the back-reflection Laue technique.&apos; The X-ray spots usually indicated a slight lineage structure of not more than a degree or two. The crystallography of slip is discussed in Schmid and Boas.&apos; The terms classical slip, cross-slip, and conjugate slip are used here with the same meanings as in ref. 3. Compression Specimen 17 (Fig. 1) was selected because the resolved shearing stresses on eight slip systems dif- fered at most by a factor of 1.2. It was hoped that in this situation such effects as cross-slip might be made more prominent. Two intersecting planes parallel to the stress axis were ground on the crystal, using optical abrasive on a lead lap. The large plane was approximately perpendicular to the plane of the axis and the primary slip direction. Its exact orientation was determined by X-ray. The smaller made an angle of 49" with the larger, measured by reflecting a small light source. The specimen was etched with nitric acid until its diameter had been reduced 0.010 in., and then was given a metallographic polish and very

Jan 1, 1954

By C. H. G. Bushell

The author maintains that selectivity in flotation depends in part on differences in rates of flotation. The rate of flotation should be directly proportional to the frequency of particle-bubble collisions. In this paper, it is suggested that floatable particles of the mineral must be distinguished from nonfloatable particles. Success in flotation is entirely a matter of selectivity. All minerals can be made to float. The problem is to make one mineral float and leave the others in the tails. For a typical ore, with all minerals naturally hydro-philic, a collector which is adsorbed by only one mineral and not at all by the others would give perfect separations. The hydrophobic, collector-coated mineral would separate cleanly from the remaining minerals. In practice, no such highly specific reagents are known. All standard collectors adsorb on most minerals to some extent, and selectivity then becomes a matter of extent of adsorption rather than adsorption on some minerals and not on others. Flotation is therefore a compromise process: if enough reagent is used to float all the desired mineral, the concentrate will be low grade; conversely, high grade concentrates can be attained only at low recovery. With many minerals coated with collector to varying degrees, selectivity appears to depend in part on each mineral floating at a different rate. The study of flotation kinetics is therefore a basic requirement to quantitative understanding of the process. REVIEW OF PREVIOUS THEORIES Before a particle enters the froth it must strike a bubble and adhere to it. Flotation rate will therefore depend on the number of particle-bubble collisions

Jan 1, 1962

By W. A., Wasley

THE EL TIGRE MINE is a highgrade silver and gold producer located in the northern part of Sonora, Mexico. It has been worked continuously since 1903, producing 50,000.000 oz. of silver and returning handsome dividends. During 1930 and 1931 falling silver prices compelled the Tigre company to make representations before the Department of Labor in Mexico City showing the necessity for drastic reductions in wages and personnel. The delays in obtaining permission from the Federal Labor Board to make these readjustments caused the company to suffer considerable financial loss during those years. After the reduction in operations was finally effected, the Sonora state government headed by Señor Francisco Elias (former Governor, now Secretary of Agriculture) and Señor Rodolfo Elias Calles (present Governor and eldest son of Señor Plutarco Elias Calles, recently Secretary of War and former President of Mexico) have cooperated with the company to make continued operation possible by revising downward certain state taxes that would have been a factor in forcing the closing of this mine. Also, permission to change our workmen over to another system of mining was essential to continued operation.

Jan 1, 1933

By V. Lessels, D. J. Buckwalter, D. H. Rose

A cyanide leaching process developed at Michigan Technological University was tested in a 12 to 24 tpd pilot plant. Flowsheet layout and equipment are described. Cyclic operation of leaching, copper precipitation by acidification, and regeneration of the cyanide leaching solution are discussed. Copper extraction and reagent consumption were as predicted by laboratory tests, and the process was proven to be technically feasible. The White Pine Copper Co., a wholly-owned subsidiary of the Copper Range Co., produces copper in the Upper Peninsula of Michigan. Production in 1965 amounted to 138 million pounds of metal. An expansion program now under way will increase this by approximately one half. The highly mechanized room and pillar mine produces 18,000 tons of ore per day, which is treated in a flotation plant. The concentrate is smelted in a reverberatory furnace, the matte is blown in converters to blister copper, and the blister copper is further refined to a uniformly high-quality Lake copper, which is 100% conductive and has a silver content of 25 to 40 oz per ton. The run of the mine ore averages 22.5 lbs of copper per ton with the ore mineral occurring in the form of chalcocite. The nature of the mineralization is such that the sand tailings average a little over 4 lbs of copper per ton. Intensive research has improved copper recovery in the flotation operation to some extent, but at the scale of present operation about 12,000 tons of sands containing over 48,000 lbs of copper go to the tailings pond daily. When the plant expansion is completed, the loss of copper will be somewhat greater. In 1959, the White Pine Copper CO, initiated a study of the recovery of copper in the sand tailings. Inasmuch as the flotation operation was considered to be optimized, a logical approach seemed to be leaching. For various reasons ammonia, ferric chloride, acid, and biological leaching were ruled out as impracticable. This led to a consideration of cyanide leach-

Jan 1, 1968

By H. H. Smith, G. J. Danek, M. R. Achter

Nickel, Type 316 stainless steel, and copper show decreases in reversed bending fatigue strength with increases in air and oxygen pressure. The leveling off of the log N us log p curve for nickel at 300" and 816°C is discussed in terms of a model in which surface saturation gives rise to a critical pressure. A metallographic study indicated that the effect of environment on crack initiation and growth rates is a function of temperature. It is shown that small amounts of impurities in inert gases may substantially reduce the fatigue strength of nickel. 1n a previous investigation1 of high-temperature bending fatigue in air and in vacuum, reversals in the effect of environment with changes in test conditions were atrributed to a mechanism consisting of two processes acting in competition. One of them, oxidation strengthening, has been demonstrated in a straightforward manner by measuring the rupture strength of a sintered nickel-oxide bond.&apos; The detailed operation of the other process, facilitation of crack propagation by surface absorption of oxygen, is uncertain. Thompson et UZ advanced the possibility that the formation of an oxide layer prevented rewelding of cracks during the compressive part of the cycle, while Wads worth and Hutchings4 further proposed that the reaction of oxygen with atoms at the tip of the crack helped the metal to break there, a possibility which has also been advanced to explain crack formation in creep.5 A difficulty in arriving at a mechanism for this lowering of fatigue strength in air is the inconsistency of some of the data. For example, Wadsworth and Hutchings4 found an approximately linear decrease of log cycles-to-fracture (N) with an increase of log pressure-of-air (p) for copper and aluminum at room temperature, while snowden,&apos; although finding a similar linear dependence for aluminum, observed a stepped curve for lead. Between l0-&apos; and 10"1 torr there is a sharp decrease in the fatigue life of lead with increasing air pressure, but plateaus, signifying no dependence on pressure, are found at higher and lower pressures. If, as discussed elsewhere: these environmental effects arise from the adsorption of gas on crack surfaces, it would be expected that as the pressure is increased a saturation point would be observed, beyond which there is no further effect. To gain further information about the form of the log N vs logp curve, it was decided in the present investigation to repeat this type of measurement at elevated temperature on nickel, copper, and Type 316 stainless steel. An additional object of this program was

Jan 1, 1963

By David White

The submission of carefully prepared estimates of the oil reserves of the United States calls for no apology or explanation. In this country, petroleum is a rapidly wasting asset and an occasional appraisal of the amount remaining in the ground is a simple business procedure to safeguard the general welfare and the prosperity of the republic. The rumored danger of failure, in the near future, of the foreign supply on which we depend in constantly increasing amounts makes the immcdiatc review of our available reserves the more imperative. The Director of the U. S. Geological Survey, in March, 1921, requested the American Association of Petroleum Geologists to cooperate in the preparation of new estimates of the oil remaining in the graund. With the members of this cooperative committee, state geologists, regional geologists, oonsulting specialists, and geologists and engineers attached to many companies worked both independently and through subcommittees. The reports for districts, counties, fields, etc., after review by local or state committees, were discussed by the joint cooperative committee in conferences that, in some instances, were attended by the original compilers of estimates for the states or smaller areas. Finally the data were reviewed and the figures revised by the joint cooperative committee, so that the totals given in Table 1 represent the opinion of the committee, which was composed of F. W. DeWolf, W. E. Wrather, Roswell H. Johnsont Wallace A. Pratt, Alexander W. McCoy, Carl H. Beal, C. T. Lupton, G. C. Matson, K. C. Heald, W. T. Thom, Jr., A. E. Fath, Kirtley F. Mather, R. C. Moore, and David White, chairman. Oil geologists and engineers understand well the speculative nature and the chances of error that must attend estimates of the petroleum in the ground when they cover not merely prospective territory, but possible territory. Neither the character nor the magnitude of the factors of uncertainty is to be disguised. Further, the inevitable differences of opinion between geologists of equal rank and experience must also bc

Jan 1, 1923

By David White

The submission of carefully prepared estimates of the oil reserves of the United States calls for no apology or explanation. In this country, petroleum is a rapidly wasting asset and an occasional appraisal of the amount remaining in the ground is a simple business procedure to safeguard the general welfare and the prosperity of the republic. The rumored danger of failure, in the near future, of the foreign supply on which we depend in constantly increasing amounts makes the immcdiatc review of our available reserves the more imperative. The Director of the U. S. Geological Survey, in March, 1921, requested the American Association of Petroleum Geologists to cooperate in the preparation of new estimates of the oil remaining in the graund. With the members of this cooperative committee, state geologists, regional geologists, oonsulting specialists, and geologists and engineers attached to many companies worked both independently and through subcommittees. The reports for districts, counties, fields, etc., after review by local or state committees, were discussed by the joint cooperative committee in conferences that, in some instances, were attended by the original compilers of estimates for the states or smaller areas. Finally the data were reviewed and the figures revised by the joint cooperative committee, so that the totals given in Table 1 represent the opinion of the committee, which was composed of F. W. DeWolf, W. E. Wrather, Roswell H. Johnsont Wallace A. Pratt, Alexander W. McCoy, Carl H. Beal, C. T. Lupton, G. C. Matson, K. C. Heald, W. T. Thom, Jr., A. E. Fath, Kirtley F. Mather, R. C. Moore, and David White, chairman. Oil geologists and engineers understand well the speculative nature and the chances of error that must attend estimates of the petroleum in the ground when they cover not merely prospective territory, but possible territory. Neither the character nor the magnitude of the factors of uncertainty is to be disguised. Further, the inevitable differences of opinion between geologists of equal rank and experience must also bc

Jan 1, 1923

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

This paper summarizes investigations during 1949 on three pilot plants for drying low-rank fine coal by entrainment in hot gases. Detailed operating results on processing seven coals having moisture ranging from 24 to 62 pct, formulas and data for calculating performance of large units, and general conclusions on the problem of drying fine coal are presented.

Jan 9, 1950

By P. J. Potter

PRACTICALLY all tin-base babbitt metals used in engine bearings are made to customers&apos; specifications, which are many and varied. The copper ranges from 3 to 8 per cent. and the antimony from 4 to 13 per cent.; generally, the babbitt with lower copper content will contain from 4 to 8 per cent. antimony and that with higher copper will have from 7 to 13 per cent. antimony. The allowable lead content varies from 0.20 to 2.00 per cent. If a solder is used as a bonding material instead of tin, the resulting material in the finished bearing will have a higher lead content, but it would not be enough to bring the lead above specificational limits. Impurities such as iron, arsenic and bismuth should be determined in the tin before using to insure a uniform and high-grade product. These points must be considered when segregating and grading the waste that comes from the finishing operations on lined bearings.

Jan 1, 1929

By R. E. Hobart

The development of the Hauto power plant and the claims made by various engineers that electricity was more economical than steam for power purposes in and about the mines, led the Lehigh Coal and Navigation Co., in 1911, to conduct a test to ascertain the consumption of power used by a large steam hoisting engine. The question being one on which no reliable information could be found, and the opinion of various engineers differing to such an extent, it was decided that a test under actual operating conditions was necessary. This was arranged for at one of the collieries. The engine selected was 2 30 by 60-in. (76.2 by 152.4-cm.) piston-valve engine of modern type, and comparatively new. The boiler plant from which the hoist received its power was about 600 ft. (182 m.) from the engine, the latter being fed by a 10-in. (25.4-cm.) steam line insulated with magnesia pipe covering. Two batteries of the boiler plant, aggregating 1200 hp., were cut off and fed into a separate steam line leading direct to the hoist engine. Steam auxiliaries, consisting of feed-water pump and blowers, were fed by the boilers in test, and their consumption charged against the hoist. A barrel weighing device was installed, as it was felt that this would be the most accurate way of determining the consumption of water. The fuel, No. 3 buckwheat, was carefully weighed, and every precaution was taken to make the test accurate in every particular. The test was run for 1 week, or a total of 168 hr. A record was kept of the number of trips hoisted or lowered, and continuous indicator cards were taken. One particular set of cards was taken with the hoist operating balanced, and also with the hoist operating with no counterweight other than the empty cage, the coal in the car being weighed. The speed of the hoist was taken by means of a graphic recording instrument which registered the number of revolutions every 5 sec. From this record, speed-time curves were plotted.

Jan 1, 1918

[NORTH AMERICA UNITED STATES ALABAMA America.-Dyar, 0. C. Anniston.-Klugh, B. 0. Ashland.-Pallister, H. D. Auburn.-*Poor, R. S. Bessemer.-Creveling, J. 0. Lindstrom, W. A. McKenzie, W. 0. Mitchell, F. R. Birmingham.-Abbott, C. E. Aldrich, T. H., Jr. Armstrong, 0. R. Beard, A. H. Benjamin, F. S. Blackburn, E. 0. Blair, A. J. Blair, 0. S. Blair. H. W.]

Jan 1, 1946

By F. L. Robertson, C. H. Bacon, J. W. Till

A description is given of the development of a process for making low sulphur steel on furnaces fired with 2 1/2 pet sulphur oil. Slag and metal are analyzed at melt. Slag weight is estimated. A simple method for calculating necessary lime additions for making 0.025 and 0.035 pct sulphur steel is given. DURING the last 15 years John Summers & Sons have been confronted increasingly by the problem of producing a low sulphur steel from high sulphur materials. The principles adopted, the processes used, the data collected on a furnace specially set aside to study this problem, and speculation on future developments are dealt with in this paper. This study is limited to two melting shops which make sheet steel only, of the following analysis: C, 0.06 pct; Mn, 0.33; S, 0.04, 0.035, 0.03, and 0.025; and P, 0.02 and 0.015. All the processes described have reference to the manufacturing of this steel and are not necessarily applicable, without modification, to other steels. The raw materials used are given in Table I. Development of Process About 1934, Schenck&apos;s work on phosphorus was applied to frame quantitative rules for the removal of phosphorus, with great success. We could however, neither in our own work nor in the voluminous literature, find any assistance to frame quantitative furnace rules to remove sulphur. Perhaps, we should have persevered with Schenck&apos;s outstanding chapter on sulphur in The Physical Chemistry of Steelmaking, but we missed the importance of one (Ss) sentence on p. 478. "Evidently this ratio -------- is [S] the important quantity to be considered from the point of view of chemical reaction; it is clear that the sulphur content of the steel alone, as frequently given in the literature cannot suffice for judging the success of the reaction unless special details are given regarding the sulphur charge and the amount of slag."&apos; We were not however very successful in getting values for Schenck&apos;s "n" that agreed with (S) arrived at by our direct analysis. [S] In addition to Schenck&apos;s work much attention was given to V value. The formula selected for V was: (SCaO) — 1.57 (2 P2O5) (S SiO2) After plotting many charges a curve was obtained, based upon the relationship between the percentage sulphur in the finished steel and V value, Fig. 1. For a long time this was used as standard with about a 60 to 70 pct success on producer-gas fired furnaces, if all uprising scrap was used and no bought scrap used. An inspection of this curve shows clearly that it is not a quantitative manufacturing process, and that it was not sure enough to go to the trouble and expense of doing stage analysis for P,O,,

Jan 1, 1952

By B. M. Bird

DURING the past four years the experimental work in coal washing carried on by the U. S. Bureau of Mines and the University of Washington has been devoted mainly to the development of special methods for the tabling of fine sizes of coals that are difficult to wash. One important outgrowth of this investigation has been the application of hindered-settling classification to coal tabling. Although used for many years in ore dressing, the term "classification" as used here may be new to many readers. It may be defined as the separation obtained when a bed of particles is held in a uniform state of suspension in an upward current of water, either continuous or pulsating, just sufficiently strong to keep the bed in a mobile condition. A common example of this type of separation is a jig in which there is no suction stroke. When the hindered-settling classifier is used with the coal-washing table, the fine-size raw coal, instead of being fed direct to tables, as is the prevailing practice, is first fed to the classifier and grouped, according to the rate of settling against the upward current of water, into perhaps six separate products. After dewatering, each of these except the sixth, a refuse product, is elevated to a seperate feed bin and washed on a separate table. This paper gives a brief discussion of the basis for the use of hindered-settling classification of a table feed; a description of the experimental work; a comparison of the results obtained with and without classification; and a discussion of the application of the classifier in commercial coal washing.

Jan 1, 1928

By C. A. Rowland, M. D. Flavel

There are seven basic comminution circuits that can be used to prepare beneficiation feed from the product of a primary crusher. This paper gives a comparison of these circuits based on average process and economical data. The paper is principally devoted to a discussion of the test work required to obtain the process data needed to evaluate each circuit. The results of the test work in conjunction with geographic and economic factors can be used to narrow down the choice of possible circuits to one, two, or three before making a detailed engineering study.

Jan 1, 1985