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By Richard W. Bayley

The Atlantic City district encompasses several districts and has been previously called by different names, e.g., Atlantic gold district, Atlantic City-South Pass mining district, and Sweetwater mining district. The district is located in a terrane of folded and faulted early Precambrian metamorphic rocks which lies near the south end of the Wind River Range in Fremont County, Wyoming, Historically it is a gold district of little production, but today is an iron district of considerable importance. Gold was discovered in placer deposits about 1842 and in quartz-arsenopyrite veins about 1867, after which there occurred a mining boom lasting a few years. Minor, intermittent gold production from placers and lodes has continued to date, but the total production has probably not exceeded a few millions of dollars. Past lode mining was generally shallow ( in the oxidized zone), the deepest about 3 60 feet. The deeper, less oxidized and sulfide ores on the same lodes have not been explored. They appear to hold the only hope for future gold mining. Sedimentary iron-formation, similar to the Lake Superior taconite, occurs in the northern part of the district. Locally, where the ironformation is thickened by folds and faults, sufficient tonnages are present to be minable, as at Iron Mountain, where the U.S. Steel Corporation produces about 1.3 million tons of iron-ore pellets annually. Exploration of the iron deposits began in 1954, and the first ironore pellets were shipped to Provo, Utah in August 1962. Indicated reserves of iron-formation (about 30 per cent Fe) are reported to be about 300 million tons.

Jan 1, 1968

By A. D. Zunkel, A. H. Larson

Equilibrium arsenic contents of Pb-As alloys in contact with PbO-As2O3 slags containing less than 30 mol pct As2O3, were determined at 650°, 700: and 750 C in an inert at?rzosphere. In this temperature range, the arsenic content of the alloy increased with increasing temperature in the single-phase liquid slag region and decreased with increasing temperature in the two-phase slag region and the single-phase solid-solution slag region. The PbO-As2O3 phase diagram below 22 mol pct As2O3 was determined by thertrzal analysis and by application of a log ?As2O3/?PbO vS log XAs2o3 /x3pbO plot determined from the equilibrium actiuity data. The resulting phase diagram was not well-defined since the eutectic temperature was not detected in the thermal analysis experiments, although a region of terminal solid solubility of As2O3 was found. Results from the phase diagram determination are compared with an existing diagram in the literature. THIS experimental investigation is an extension of a study by the authors1 on the slag-metal equilibria in the systems concerned with commercial lead refining processes such as softening and dross fuming. The first part of this investigation was a study of the slag-metal equilibria in the Pb-PbO-Sb2O3 system. The only experimental work previously done on the Pb-PbO-As2O3 system was by Pelzel2,3 in which the phase diagram for the PbO-As2O3 system was determined below 50 wt pct As2O3 and the equilibrium constant for the reaction 3Pb + As2O3 + 3PbO + 2As was determined as a function of temperature. No slag-metal equilibrium data have been determined. It is due to the scarcity of information regarding the Pb-PbO-As203 system that this work was undertaken. This paper describes the determination of the slag-metal equilibria in the Pb-PbO-As203 system by equilibrating Pb-As alloys with PbO-As2O3 slags in an inert atmosphere, the effect of 1 wt pct additions of bismuth and copper on the slag-metal equilibria, and the PbO-As2O3 phase diagram both by thermal analysis and the use of the slag-metal equilibria data. EXPERIMENTAL Materials. The materials used in this investigation were analytical reagent-grade and assayed as follows: 1) 99.8 pct PbO (0.014 pct insoluble in CHJCOOH, 0.02 pct not precipitated by H2S, 0.1 pct CaO, and 0.08 pct SiO2); 2) 99.95 pct As2O3; 3) 99.99 pct Pb; 4) 99.0 pct As; 5) 99.99 pct Cu; and 6) 99.97 pct Bi. Room-temperature X-ray patterns revealed no detectable impurities in any of these materials. Apparatus for Equilibrium and Thermal-Analysis Determinations. The resistance-heated crucible furnace used in this investigation employed nichrome elements and was mounted so that it could be raised to surround the reaction tube during each experiment and, subsequently, lowered. A schematic diagram of the apparatus is shown in Fig. 1. Each charge was heated in a 3+-in.-OD by 31/2-in.-high 416 stainless-steel crucible placed in a 41/4-in.-1D by 18-in.-long fused-silica reaction tube which was closed at one end. On a shoulder around the crucible was placed a 3: -in.-OD by 12-in.-long open-end fused-silica condenser tube. The open end of the reaction tube was covered by a water-cooled brass cap with ports for 1) admitting an inert atmosphere to the system through a stopcock, 2) introducing a stainless-steel, motor-driven, paddle stirrer into the crucible, 3) evacuating the system with a mechanical vacuum pump, and 4) sampling the melt with Vycor sampling tubes. The brass cap was fitted to the open end of the reaction tube with a silicone gasket and collar clamp. The furnace temperature was controlled by a Barber-Coleman Capacitrol controller and a chromel-alumel thermocouple. Due to the corrosiveness of the melt, the controlling thermocouple also served as the measuring thermocouple. The temperature of the melt was calibrated against the controller temperature and was checked periodically during each test with a Vycor-enclosed calibrated chromel-alumel thermocouple. The temperature measurement and control can be considered accurate to ±3°C. Procedure. The charge placed in the crucible for each experiment consisted of 1000 g of Pb-As alloy and 300 g of PbO-As2O3 slag. The crucible was then placed in the reaction tube, the condenser tube was placed on the shoulder of the crucible, the silicone gasket and brass cap were fitted on the open end of the reaction tube, and the entire system was evacuated and filled with argon ten times. After the last flushing, a positive argon pressure of 1 psig was impressed on the system. The furnace was then raised to sur-

Jan 1, 1968

By D. T. Novick, M. B. Bever, B. Roessler

Ordered and disordered specimens of the alloy CusAu were cold worked and then annealed at 2880, 310°, and 340°C, together with disordered unworked specimens. me resistivity decreased during annealing mainly as a result of ordering; the resistivity measurements suggest that in cold-worked specimens the growth of antiphase domains is impeded. The strength properties increased during annealing at 288 "C; in cold-worked specimens this increase was lo a level above that for cold-worked specimens which had not been annealed. X-ray diffraction showed that after annealing for a few hours at 288ºC, the long-range order increased to a constant value; additional annealing at 288ºC resulted in the growth of antiphase domains in which the degree of order was constant. Changes in peak breadth and particle size confirmed that in cold-worked specimens the major effects of cold work were not eliminated during annealing at 288 "C, although a slight recovery process occurred. When a specimen of the alloy Cu3Au is heated to a temperature below the critical temperature (388ºC), long-range ordering tends to occur in a specimen containing less than the equilibrium amount of order characteristic of that temperature Ideally, a specimen of Cu3Au contains no long-range order if it has been quenched from above the critical temperature. Long-range order is reduced by cold work. If cold-worked specimens are heated, restoration processes such as recovery and recrystallization may occur,

Jan 1, 1963

By C. P. Proctor, J. B. Morrow

With numerous plans under consideration for coal classification, and with the advent of the Bituminous Coal Code, the intelligent sampling of coal has become increasingly important. To us it is rather significant that there have been very few fundamental studies of coal sampling in American technical literature since those of E. G. Bailey, which appeared in the Journal of Industrial and Engineering Chemistry in March, 1909. He covered the subject up to that time so thoroughly and in such an excellent manner that perhaps many other later students have felt that they could add very little to his findings. The recent studies of the British Engineering Standards Association have indicated the possibility of a somewhat different interpretation of Bailey's results, permitting the taking of smaller samples, and perhaps it will now be helpful to reopen this question. C. W. H. Holmes, in an informative series of articles on coal sampling recently published in Colliery Engineering, makes two rather interesting statements, as follows: It is scarcely surprising to find a lack of interest and knowledge regarding the sampling of coal not only in the mining industry itself but also among consumers of industrial fuel: at its best coal sampling is a dirty business, regarded as a nuisance and an expense by the commercial man, and usually as a nuisance and nothing else by the chemist. The writer can conceive only two classes of people who have never experienced any difficulty in obtaining representative samples of coal from wagons; those who have never tried to, and those who have never had the necessity or desire to check the accuracy of their work. The primary purpose of this paper is to present some tabulated data of the more significant results of a sampling investigation undertaken during the last four years to obtain certain fundamental data on the variables connected with coal sampling and to point out that often assumptions have been made that are not in accordance with the facts. One of the standard sampling methods states:

Jan 1, 1936

By C. P. Proctor, J. B. Morrow

With numerous plans under consideration for coal classification, and with the advent of the Bituminous Coal Code, the intelligent sampling of coal has become increasingly important. To us it is rather significant that there have been very few fundamental studies of coal sampling in American technical literature since those of E. G. Bailey, which appeared in the Journal of Industrial and Engineering Chemistry in March, 1909. He covered the subject up to that time so thoroughly and in such an excellent manner that perhaps many other later students have felt that they could add very little to his findings. The recent studies of the British Engineering Standards Association have indicated the possibility of a somewhat different interpretation of Bailey's results, permitting the taking of smaller samples, and perhaps it will now be helpful to reopen this question. C. W. H. Holmes, in an informative series of articles on coal sampling recently published in Colliery Engineering, makes two rather interesting statements, as follows: It is scarcely surprising to find a lack of interest and knowledge regarding the sampling of coal not only in the mining industry itself but also among consumers of industrial fuel: at its best coal sampling is a dirty business, regarded as a nuisance and an expense by the commercial man, and usually as a nuisance and nothing else by the chemist. The writer can conceive only two classes of people who have never experienced any difficulty in obtaining representative samples of coal from wagons; those who have never tried to, and those who have never had the necessity or desire to check the accuracy of their work. The primary purpose of this paper is to present some tabulated data of the more significant results of a sampling investigation undertaken during the last four years to obtain certain fundamental data on the variables connected with coal sampling and to point out that often assumptions have been made that are not in accordance with the facts. One of the standard sampling methods states:

Jan 1, 1936

By Terkel Rosenqvist

Few subjects in iron metallurgy and steelmaking have caused so much discussion and have resulted in so many divergent and partly contradictory statements as the theory of the distribution of sulphur between molten iron and slag. An interesting survey of the different points of view is given by Grant and Chipman.1 Grant and Chipman themselves conclude in their paper that, at least in the open hearth and electric steel furnaces, the governing reaction is a simple distribution like: FeS steel ? FeS alas [Ia] and MnS steel ? MnS alas [Ib] They conclude that no significant reaction takes place according to the more classical scheme: FeS steel + CaO slag ? FeOslag + CaS slag [2a] This is the reaction which is regarded by most investigators as the primary sulphur reaction in the blast furnaces, where it is accompanied by the reaction with carbon: FeO slag + C ? Fe + CO [2b] The writer suggests an alternative to the reactions mentioned above. It is the opinion of the writer that the a priori assumption of defined molecular compounds being present in molten steel as well as in molten slag, has retarded rather than promoted a solution of the sulphur problem. As for the constituents of molten steel, it is taking too much for granted to mention the presence of FeS and MnS. If this infers that the melt contains diatomic molecules, the statement must be made that at least such molecules do not exist in the solid state at room temperature. Solid FeS crystallizes in the nickel-arsenide structure, and MnS (high temperature form) in the sodium-chloride structure. In these structures the sulphur is surrounded by six metal neighbors, and no molecular groups can be isolated. When sulphur is dissolved in molten steel, it is expected to maintain the same coordination number as in the solid state. We know that sulphur in steel has a certain chemical potential and a chemical activity which can be expressed by the reaction: S steel + H2 ? H2S where the activity A. is proportional to the ratio PH2S/Ph2. We know also that this activity is influenced by the presence of other constituents. For instance, there is reason to assume that carbon and silicon increase the sulphur activity and manganese decreases it. However, the exact relationship here is scarcely known, and further investigations are needed. Concerning the molten slags, if anything should be assumed a priori it would be that they are composed of ionic rather than molecular parts. There are indications of this in that solid silicates form ionic structures, and molten slags show high electric conductivity and can be electrolyzed.2 It has been claimed by Taylor and Chipman3 that their activity investigations

Jan 1, 1949

By Terkel Rosenqvist

Few subjects in iron metallurgy and steelmaking have caused so much discussion and have resulted in so many divergent and partly contradictory statements as the theory of the distribution of sulphur between molten iron and slag. An interesting survey of the different points of view is given by Grant and Chipman.1 Grant and Chipman themselves conclude in their paper that, at least in the open hearth and electric steel furnaces, the governing reaction is a simple distribution like: FeS steel ? FeS alas [Ia] and MnS steel ? MnS alas [Ib] They conclude that no significant reaction takes place according to the more classical scheme: FeS steel + CaO slag ? FeOslag + CaS slag [2a] This is the reaction which is regarded by most investigators as the primary sulphur reaction in the blast furnaces, where it is accompanied by the reaction with carbon: FeO slag + C ? Fe + CO [2b] The writer suggests an alternative to the reactions mentioned above. It is the opinion of the writer that the a priori assumption of defined molecular compounds being present in molten steel as well as in molten slag, has retarded rather than promoted a solution of the sulphur problem. As for the constituents of molten steel, it is taking too much for granted to mention the presence of FeS and MnS. If this infers that the melt contains diatomic molecules, the statement must be made that at least such molecules do not exist in the solid state at room temperature. Solid FeS crystallizes in the nickel-arsenide structure, and MnS (high temperature form) in the sodium-chloride structure. In these structures the sulphur is surrounded by six metal neighbors, and no molecular groups can be isolated. When sulphur is dissolved in molten steel, it is expected to maintain the same coordination number as in the solid state. We know that sulphur in steel has a certain chemical potential and a chemical activity which can be expressed by the reaction: S steel + H2 ? H2S where the activity A. is proportional to the ratio PH2S/Ph2. We know also that this activity is influenced by the presence of other constituents. For instance, there is reason to assume that carbon and silicon increase the sulphur activity and manganese decreases it. However, the exact relationship here is scarcely known, and further investigations are needed. Concerning the molten slags, if anything should be assumed a priori it would be that they are composed of ionic rather than molecular parts. There are indications of this in that solid silicates form ionic structures, and molten slags show high electric conductivity and can be electrolyzed.2 It has been claimed by Taylor and Chipman3 that their activity investigations

Jan 1, 1949

By George E. Beavers

Under the title used for this paper, the converter practice of the Tennessee Copper Co. is described in the TRansactions of the Institute.l That article contains the following statement: "The limit in capacity of the Peirce-Smith converter in terms of concentrate has not been reached but approximately as high as one ton of concentrate has been handled for each ton of liquid matte charged and good working conditions have been maintained." It is not the intention of the writer to review the conditions leading up to the introduction of a Peirce-Smith converter at Copperhill, but, instead, to offer later data to show how satisfactory the installation has been. Converter Campaign, June 15, 1930 to August 23, 1931 Blowing time (converter days)............... 349.90 Number of blows........................... 427.00 Tons matte charged........................ 38,335.00 Copper assay of matte, per cent.............. 13.67 Tons blister copper made.................... 9,246.00 Time to blow one ton of copper, minutes...... 54.50 Air used per minute, cu. ft................... 12,500.00 Blast pressure, lb........................... 12.80 Air per ton copper, cu. ft.................... 681,330.00 Air per ton iron, cu. ft....................... 199,040.00 Tons flux used............................. 11,845.00 Tons flux per ton matte..................... 0.31 Tons flux per ton copper.................... 1.28 Tons flux per ton iron....................... 0.37 Tons concentrate treated.................... 28,960.00 Copper assay of concentrate, per cent......... 17.13 Tons concentrate per ton matte.............. 0.76 Tons material charged per converter day...... 235.70

Jan 1, 1934

By Heinrich Freyn

THE American iron and steel manufacturer finds himself to-clay barely at the threshold of enormous possibilities for practicing rational economy in the use of fuels. The fuel cost is by no means the smallest of items which enter into the cost of manufacture of iron and steel products, but the great wealth of this country in high-grade fuels, together with a protective tariff, has heretofore made fuel economy relatively unimportant. In Germany, the use of raw coal for the purposes of iron and steel manufacture has been all but abandoned and large sums of money, formerly irrevocably wasted, are now being saved. When scanning the literature of this country devoted to the application of gas power, but little information will be found on a subject which, however, is destined to become of considerable importance in the United States, namely, the use of coke-oven gas for the generation of power in gas engines. This lack of information can be traced to the fact that the regenerative by-product coke oven with its large amount of surplus gas is still far from having attained in this country the universal application which it has found in Europe. This statement needs hardy any corroboration other than the evidence of statistical investigations made by E. W. Parker of the United States Geological Survey, according to which only approximately 25 per cent. of the total coke produced in this country in 1912 was made in by-product coke ovens, whereas 75 per cent. was produced in bee-hive ovens. This entailed the enormous loss of over $80,000,000 in non-recovered by-products, such as gas, tar, and ammonia, without considering the substantial loss due to the non-recovery of benzol from the coal. There are two more or less distinct fields of application of the by-product coke oven: one, to furnish as the main product metallurgical coke of specified quality, and surplus coke-oven gas as the by-product, and the other, to carbonize high-volatile coals, and thus produce as the main product large quantities of gas, high in heat value and rich in illuminants, adapted primarily for domestic heating and lighting, while domestic coke becomes the by-product.

Jan 4, 1914

By J. D. Rice, P. L. Wearden, W. L. Hensou

Solutions to the unsteady-state partial water-drive reservoir performance problem can be obtained through the use of analogue computers or high-speed electronic digital computers. The solutions that have previously been resolved for use on digital computers, however, demand a knowledge of the aquifer parameters. Generally, the analogue computers now in use do not require this knowledge. A solution is presented herein where the aquifer performance, expressed in terms of difference equations, is related to the reservoir performance as expressed by the modified Scbiltbuis material-balance equation. A numerical procedure for a medium-sized digital computer also is presented in which a solution to the set of equations defining the aquifer and reservoir performance is obtained and the aquifer parameters (permeability and sand thickness) are automatically optimized while simultaneously matching the known pressure behavior. Predicted pressure behavior can be calculated using rates from any assumed future production practice. The procedure provides an output format which presents the cumulative, incremental and average rate of natural water influx, the per cent gas-cap expansion, the calculated reservoir pressure, the measured reservoir pressure, and the difference between the measured and calculated pressures. Results of a test problem are presented in comparison with results obtained by the Bruce Analyzer, Ohio Oil Co. 's Pace General Purpose Analogue Computer, and Sun Oil Co.'s Single-Pool Electronic Reservoir Analyzer. These results indicate that unsteady-state reservoir perlormance for a single-pool system can be adequately simulated by a numerical method employing a digital computer and that the special-purpose analogue computer can be supplanted by this method. INTRODUCTION The numerical approach to the unsteady-state partial water-drive reservoir performance problem heretofore has been limited in that solutions have been resolved only for certain fixed parameters. Also, prior to the advent of high-speed electronic digital computers in the last few years, calculations were so tedious and time-consuming that they ordinarily were neglected, or some other means of obtaining the solutions were used. The analogue approach had been used for obtaining solutions to the transmission-of-heat problem and, consequently, means were developed for applying the electrical analogue to the unsteady-state reservoir performance problem. Bruce submitted a paper for AIME publication in Sept. 1942, in which he described an electronic device (now called the Bruce Analyzer) for analyzing the water-drive performance of any reservoir, based on either steady or unsteady-state flow characteristics. 1 Sun Oil Co. adopted the analogue approach in 1949 and now has two special-purpose high-speed electronic reservoir analyzers in use.2,3 Two main features of Sun's analyzers are (1) the high repetitive rate of reproducing the entire history of reservoir performance cyclically a number of times per second, and (2) the elimination of the necessity of knowing the properties of the associated aquifer. Since the advent of larger and faster digital computers, the possibility of incorporating this second feature into a numerical approach appeared feasible. Previous work has been done and material published covering the application of numerical methods to the behavior of the pressure with respect to time at any point in an associated aquifer? However, the aquifer parameters must be known from some source. The material presented in Ref. 4 has been used to a great degree in developing the numerical method described herein. DIFFERENTIAL EQUATION AND BOUNDARY CONDITIONS The problem of simulating the performance of a reservoir and the associated aquifer is to determine the characteristics of the aquifer such that the

By R. Priestner, R. F. Krause

Solid, cylindrical crystals of a iron have been fractured at 10°K by the simultaneous application of tension and torsion stresses. Thus, a complex distribution of normal (sN) and shear (ss) stresses existed in a cleavage plane at fracture. In every crystal the clearage-plane normal was inclined to the specimen axis, so that a unique point, M*, in the operative cleavage plane was defined where the quantities sn and s2 N + s2 S) were maximized. It is shown that the maxirrzum resolved normal stress on the cleavage plane at fracture decreased as the resolved shear stress in the cleavage plane at M* increased. Thus, a normal stress la?*! for fracture was not satisfied. A fracture law which takes shear stress into account. and which predicts that (s2 N +s2S) should be constant, is to be preferred. However, the data is best represented by a linear dependence of sN on sS, sS which 110 physical significance can be attached. It is suggested that the fracture-nucleating -mechanism is a short-range one. A critical stress law has not yet been unequivocally established for the propagation of cracks in metals. Measurements of the tensile fracture stresses of variously oriented zinc single crystals1 failed to establish the normal stress law expressed by In Eq. [1], sN is the fracture stress, P, resolved normal to the fracture plane, and 0 is the angle between the fracture -plane normal and the stress axis. Similarly, a comprehensive investigation failed to verify the normal stress law for the tensile fracture of a-iron single crystals.' However, a subsequent investigation,3 using iron from the same ingot, presented fracture data which obeyed the normal stress law. It has been shown previously4,5 that the Griffith criterion for fracture6 does not predict such a normal stress law. This result is of particular significance in considering the fracture of metals, such as zinc and iron, which may fail by cleavage. In these metals fracture propagates in a well-defined "cleavage" plane, whether or not an applied stress is normal to it. When the cleavage-plane normal deviates from the direction of application of a uniaxial stress, for example, the crack propagates under the combined action of normal (sN) and shear (ss) stresses. When the presence of a resolved shear stress in the crack plane is taken into account, the Griffith criterion is expressed by4

Jan 1, 1967

By T. K. Prentice

The January-February number of the Journal of the Chemical Metallurgical and Mining Society of South Africa, 1943, is made up entirely of a discussion of Ball Wear in Cylindrical Mills. Following the principal paper by T. K. Prentice are several remarkably pertinent and carefully prepared discussions. Few persons in this country would be likely to see this publication, and since the information presented is new and of current interest, it seems desirable to abstract and review this publication for the benefit of the members of the A.I.M.E. The purpose of the investigation is stated in the first two paragraphs of Prentice's paper, as follows: Early in 1941, The Chamber of Mines Steel Products, Limited, initiated a series of tests under semi-working scale conditions to determine, if possible, those physical and chemical properties of a ball which make it resistant to wear when used in a cylindrical mill as a grinding medium, with the primary object of incorporating those beneficial qualities in the balls being produced at their works. This paper describes the tests carried out and discusses the results obtained. From these results it was possible to formulate a new theory relating to ball wear in cylindrical mills. This theory is presented in the second portion of this paper. Practical Ball-mill Tests in a Cylindrical Mill Part I of the paper begins with a discussion of the methods used for the manu- facture of grinding balls in South Africa. The following paragraph is of particular interest. The cast steel ingots, 4¼ in. square by 3 ft. 6 in. long, are re-heated in a gas-fired sloping stationary-hearth furnace and then rolled in a six-pass continuous mill into 25/8 in. round billets. Some of these billets are then removed and cut into suitable lengths for forging into 4 in., 3¼ in. , 3 in. or 2½ in. balls by means of three Drop and two Beche hammers, but the greater number, whilst still at about 1,000, pass directly into a set of scroll rolls which produce, in an ingenious manner, finished 3 in. balls direct from the 25/8 in. billets. Except in special cases, all balls are allowed to cool in the atmosphere and are not water quenched. Details of the Tests In order to enable the reader to evaluate properly the results and conclusions in the rest of the paper and in the discussions following the paper, most of the information presented by the author under this heading is quoted: The testing unit was a nominal 30 in. diameter and 18 in. long centre discharge Denver Ball Mill. The ore pulp to the ball mill comprised a small portion of the normal feed to one of the fourteen City Deep 5 ft. 6 in. X 22 ft. tube mills, and the grinding media consisted of the balls under test plus other balls to give the desired load. The running time of the test-mill was divided into approximately equal periods of five days 13½ hours each. The usual procedure was to examine and make up the load

Jan 1, 1947

By T. K. Prentice

The January-February number of the Journal of the Chemical Metallurgical and Mining Society of South Africa, 1943, is made up entirely of a discussion of Ball Wear in Cylindrical Mills. Following the principal paper by T. K. Prentice are several remarkably pertinent and carefully prepared discussions. Few persons in this country would be likely to see this publication, and since the information presented is new and of current interest, it seems desirable to abstract and review this publication for the benefit of the members of the A.I.M.E. The purpose of the investigation is stated in the first two paragraphs of Prentice's paper, as follows: Early in 1941, The Chamber of Mines Steel Products, Limited, initiated a series of tests under semi-working scale conditions to determine, if possible, those physical and chemical properties of a ball which make it resistant to wear when used in a cylindrical mill as a grinding medium, with the primary object of incorporating those beneficial qualities in the balls being produced at their works. This paper describes the tests carried out and discusses the results obtained. From these results it was possible to formulate a new theory relating to ball wear in cylindrical mills. This theory is presented in the second portion of this paper. Practical Ball-mill Tests in a Cylindrical Mill Part I of the paper begins with a discussion of the methods used for the manu- facture of grinding balls in South Africa. The following paragraph is of particular interest. The cast steel ingots, 4¼ in. square by 3 ft. 6 in. long, are re-heated in a gas-fired sloping stationary-hearth furnace and then rolled in a six-pass continuous mill into 25/8 in. round billets. Some of these billets are then removed and cut into suitable lengths for forging into 4 in., 3¼ in. , 3 in. or 2½ in. balls by means of three Drop and two Beche hammers, but the greater number, whilst still at about 1,000, pass directly into a set of scroll rolls which produce, in an ingenious manner, finished 3 in. balls direct from the 25/8 in. billets. Except in special cases, all balls are allowed to cool in the atmosphere and are not water quenched. Details of the Tests In order to enable the reader to evaluate properly the results and conclusions in the rest of the paper and in the discussions following the paper, most of the information presented by the author under this heading is quoted: The testing unit was a nominal 30 in. diameter and 18 in. long centre discharge Denver Ball Mill. The ore pulp to the ball mill comprised a small portion of the normal feed to one of the fourteen City Deep 5 ft. 6 in. X 22 ft. tube mills, and the grinding media consisted of the balls under test plus other balls to give the desired load. The running time of the test-mill was divided into approximately equal periods of five days 13½ hours each. The usual procedure was to examine and make up the load

Jan 1, 1947

By C. D Looker

Common salt is mentioned in the most ancient writings as an important article of diet. It is fairly certain that it was used by men and animals long before the dawn of civilization. Its presence in natural springs has often been the determining factor in colonization; wars have been fought for possession of salt deposits; trade routes have been established expressly for the purpose of traffic in salt. Naturally this led men into adventures and discoveries that altered history and advanced civilization. In the early days the uses for salt were few and the methods of manufacture were crude and the purity low when compared to salt of the present day. Then the salt came largely from springs, shallow wells or the ocean, while today deep wells must be drilled to the natural brines or beds of rock salt and spacious shafts sunk so that the ancient deposits may be made to yield their supply of stored mineral. Mining and Purification The quantity of available salt seems inexhaustible. It occurs in practically every country of the globe. About 20 million tons is produced and consumed annually in the world. Of this amount about 7 million tons is produced in the United States; 13/4 million tons of rock salt, 2 1/4 million tons as evaporated salt and 31/2 million tons in brine used in the chemical industry. Mined rock salt is used in its natural state without purification. It is classified for the trade by crushing and screening. The average purity of rock salt mined in the United States is about 98.3 per cent pure sodium chloride1. The usual impurities are gypsiferous shale, anhydrite and silica, with traces of potassium, magnesium and calcium compounds. Refined or evaporated salt is made by crystallizing the salt from brine in open pans or vacuum pans. A comparatively small amount of salt is made from ocean water or the brines of inland lakes or seas, sometimes by solar evaporation. By far the larger amount of brine is obtained by drill-

Jan 1, 1938

By C. D. Looker

Common salt is mentioned in the most ancient writings as an important article of diet. It is fairly certain that it was used by men and animals long before the dawn of civilization. Its presence in natural springs has often been the determining factor in colonization; wars have been fought for possession of salt deposits; trade routes have been established expressly for the purpose of traffic in salt. Naturally this led men into adventures and discoveries that altered history and advanced civilization. In the early days the uses for salt were few and the methods of manufacture were crude and the purity low when compared to salt of the present day. Then the salt came largely from springs, shallow wells or the ocean, while today deep wells must be drilled to the natural brines or beds of rock salt and spacious shafts sunk so that the ancient deposits may be made to yield their supply of stored mineral. Mining and Purification The quantity of available salt seems inexhaustible. It occurs in practically every country of the globe. About 20 million tons is produced and consumed annually in the world. Of this amount about 7 million tons is produced in the United States; 13/4 million tons of rock salt, 2 1/4 million tons as evaporated salt and 31/2 million tons in brine used in the chemical industry. Mined rock salt is used in its natural state without purification. It is classified for the trade by crushing and screening. The average purity of rock salt mined in the United States is about 98.3 per cent pure sodium chloride1. The usual impurities are gypsiferous shale, anhydrite and silica, with traces of potassium, magnesium and calcium compounds. Refined or evaporated salt is made by crystallizing the salt from brine in open pans or vacuum pans. A comparatively small amount of salt is made from ocean water or the brines of inland lakes or seas, sometimes by solar evaporation. By far the larger amount of brine is obtained by drill-

Jan 1, 1938

By Charles Hart

The art of electric smelting came with the turn of the present century and owes its existence to the introduction of alternating current, which found its first wide use in the establishment of the great power plants at Niagara Falls, N. Y. Prior to that time current was largely direct. The furnaces were electrolytic in nature and found their first metallurgical use in the production of aluminum, and this process still prevails. The normal production of pig iron up to that time was carried on in blast furnaces in which charcoal, coke and anthracite were used as the fuel. This new electric process called for sufficient carbonaceous material to reduce the oxides and impregnate the pig iron with sufficient carbon, but the remainder of the energy was to be supplied by electric power. The chemical reactions are essentially the same in electric smelting as obtains in the older blast furnace. All of these data are so well known in both methods as to need no discussion at this time, for they can be found in scientific : publications and books written during the first two decades of this century. Types of Furnaces The 40 years devoted to the practice of electric smelting may be divided into two . periods of 20 years each. The first was devoted to experimentation with furnace types and with many publications on theories and processes. The second half of this term was devoted to perfecting the types of furnaces and their active commercial operation. The arc furnace may be found in three types: (1) the direct-heating arc furnace with the positive electrode included in the bottom and the other electrode suspended above; (2) the independent arc furnace with its electrodes entirely above the charge, which of necessity depends upon heat reflected from the roof; (3) the direct-heating series arc furnace equipped with two or more electrodes suspended from the roof. Of these three types, only the last is used in electric smelting of iron ore at the present time. The elimination of the other types automatically removed many of the early designs, to the end that there are now only two general types in operation: (I) the " Electrometalls" or high-shaft furnace of Sweden, and (2) the low-shaft or pit-type furnace as offered by Tysland-Hole and Siemens-Halske. The "Electrometalls" furnace takes its name from the corporation formed by the engineers that had undertaken its perfection. After trying several designs they brought forward the furnace shown in Fig. I, which by the way is the highest development of this type of furnace as of today.l Note the trend in the lines, which have kept pace with modern blast-furnace design. The high-shaft furnace lends itself to skip filling and is fitted with tuyeres immediately under the crucible roof, whereby the gases are recycled and the roof preserved. The failure of the Electrometalls furnace in Norway, because of its inability to meet

Jan 1, 1941

By Charles Hart

The art of electric smelting came with the turn of the present century and owes its existence to the introduction of alternating current, which found its first wide use in the establishment of the great power plants at Niagara Falls, N. Y. Prior to that time current was largely direct. The furnaces were electrolytic in nature and found their first metallurgical use in the production of aluminum, and this process still prevails. The normal production of pig iron up to that time was carried on in blast furnaces in which charcoal, coke and anthracite were used as the fuel. This new electric process called for sufficient carbonaceous material to reduce the oxides and impregnate the pig iron with sufficient carbon, but the remainder of the energy was to be supplied by electric power. The chemical reactions are essentially the same in electric smelting as obtains in the older blast furnace. All of these data are so well known in both methods as to need no discussion at this time, for they can be found in scientific : publications and books written during the first two decades of this century. Types of Furnaces The 40 years devoted to the practice of electric smelting may be divided into two . periods of 20 years each. The first was devoted to experimentation with furnace types and with many publications on theories and processes. The second half of this term was devoted to perfecting the types of furnaces and their active commercial operation. The arc furnace may be found in three types: (1) the direct-heating arc furnace with the positive electrode included in the bottom and the other electrode suspended above; (2) the independent arc furnace with its electrodes entirely above the charge, which of necessity depends upon heat reflected from the roof; (3) the direct-heating series arc furnace equipped with two or more electrodes suspended from the roof. Of these three types, only the last is used in electric smelting of iron ore at the present time. The elimination of the other types automatically removed many of the early designs, to the end that there are now only two general types in operation: (I) the " Electrometalls" or high-shaft furnace of Sweden, and (2) the low-shaft or pit-type furnace as offered by Tysland-Hole and Siemens-Halske. The "Electrometalls" furnace takes its name from the corporation formed by the engineers that had undertaken its perfection. After trying several designs they brought forward the furnace shown in Fig. I, which by the way is the highest development of this type of furnace as of today.l Note the trend in the lines, which have kept pace with modern blast-furnace design. The high-shaft furnace lends itself to skip filling and is fitted with tuyeres immediately under the crucible roof, whereby the gases are recycled and the roof preserved. The failure of the Electrometalls furnace in Norway, because of its inability to meet

Jan 1, 1941

A GIFT of $10,000 to the endowment of Engineering Foun-dation, made by the Central Hudson Gas and Electric Corporation, has been announced. In transmitting this gift T. R. Beal, president of the corporation, made the follow-ing interesting statement: "You may be interested to know that this amount was in fact earned by our employees by their more efficient opera-tion of the physical properties under their care. It rep-resents $10,000 of return premiums paid during the past year on accident insurance, as the result of the company's excellent record in preceding years, so that we take greater satisfaction in making this gift as it is related, in some degree at least, to the higher standards of engineering and operating practice to which the work of the Foundation per-tains."

Jan 4, 1928

By A. Raja, C. C. Harris

Apparatus for maintaining constant — but adjustable — pulp level and rate of froth removal are described. The results of preliminary experiments into the effect of scraping speed on the rate of solid and liquid recovery with a semi-batch system are presented. Glass spheres restricted to a narrow size range is the solid under investigation. Graphical methods for estimating the mass of recoverable material and the zero error are given. The semi-batch procedure1 is extensively used in laboratory flotation testing, though the method is subject to variability of moderate severity. Raffinot2 concluded from a statistical study of the difference between replicate tests, that the ptincipal source of errot was the test itself and not sampling or chemical analysis. These differences can be ascribed principally to changes in pulp volume and reagent concentration, and uncontrolled rate and depth of scraping. Semi-batch testing has been criticized1 on these grounds as a means of investigating flotation kinetics. This paper discusses and describes modifications to laboratory flotation equipment which reduce the number of manual operations and lead to improved control. Kinetic studies were the primary reason for developing the apparatus described. It should, however be a useful aid in normal flotation testing. Some preliminary results of experiments are presented.

Jan 1, 1967

By B. C. Raynes

IN the early part of 1952, under the auspices of the U. S. Atomic Energy Commission, Horizons Inc. undertook an investigation dealing with the preparation of high purity thorium metal in order to develop an economically competitive production process for this metal based on a recovery of the metal by a fused salt electrolysis, since a survey of the literature of thorium chemistry indicated that electrolytic methods had not received much attention, despite the consideration that electrochemical production methods often offer attractive commercial potentialities. Thorium metal has been prepared by methods which fall into four general categories: 1) reduction of halides or double halides; 2) reduction of the oxide; 3) thermal decomposition of halides; and 4) electrolytic methods. Some of the important results obtained by various investigators since about 1900 in recorded attempts to prepare pure thorium metal are summarized in Table I. A bibliography of pertinent literature is included in this paper. Thorium is strongly electropositive and possesses a great affinity for oxygen and nitrogen. As a result the prospects for the production of high purity thorium from aqueous media or oxygen bearing compounds are remote. The most promising and profitable area of investigation for production of thorium metal lies in fused salt electrolysis. Thorium has been produced previously by this technique from halide salts in fused alkali chloride eutectic mixtures. Until recently, however, the lack of known large uses for thorium has inhibited the investigation of these methods. The basic problem in an investigation into fused salt production methods for thorium is to develop a suitable fused salt system for the economic preparation of thorium. The first portion of the work in these laboratories was a small scale investigation of the electrolysis of potassium thorium fluoride, KThF5, under a variety of conditions. High purity thorium was prepared in this manner, but the process did not appear to offer the simplest and most attractive continuous commercial operation. The work reported here relates to an all-chloride system employing thorium chloride as the cell feed. ThCL4 was produced as an anhydrous material stabilized in molten sodium chloride. This product, further diluted in NaCl, was subjected to electrolysis. The electrolytic process has been studied in cells capable of holding up to 50 lb total salt charge. Ductile, pure thorium metal has been reproducibly obtained in this operation. Considerable experimentation was carried out in the preparation of ThCL as a cell feed free from moisture, oxides, and other impurities. Thorium was cathodically deposited as a coarsely crystalline powder interspersed with residual bath salts. Aqueous procedures for separating and recovering the metal were developed. The thorium metal powder obtained has been evaluated with respect to its hardness, chemical purity, and certain other physical properties. Preparation of ThC1,—Thorium chloride, and the chlorides of some other metals (aluminum, beryllium, tantalum, zirconium, etc.) cannot be obtained from aqueous solution by direct heating and evaporation. Evaporation and heating of a solution of thorium chloride yields a hydrated salt or partially hydrated salt, and finally Tho,. Anhydrous thorium chloride is obtained by forming a hydrated ammonium chloride complex salt in acidified water solution, dehydrating this compound, and finally decomposing the double salt and subliming the ammonium chloride component. Various methods for the preparation of anhydrous thorium chloride, in accordance with this concept, were investigated as indicated in Table I1 from hydrated thorium chloride (ThCl, . 3H2O), and from thorium oxycarbonate (ThOCO,). In the method found most satisfactory for this work, thorium oxycarbonate was dissolved in an excess of concentrated HC1 to give a clear yellow solution. At least 2 mol NH,Cl were added to the solution per mol of contained thorium. Ammonium thorium chloride hydrate was crystallized from solution by cooling and saturation with HC1. The resultant complex, (NH,Cl),ThCl;xH,O, was dehydrated by evaporation at 130° to 150°C and thoroughly dried at about 250°C. The dried ammanium thorium chloride was then mixed with sodium chloride in proportions to form NaThCL or to produce a ThCL4-NaCl mixture with a given thorium content. The ammonium chloride was sublimed off at temperatures above 500°C in an inert atmosphere furnace to give an anhydrous chloride source electrolyte, essentially free of Tho, and of NKC1.

Jan 1, 1958