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By S. F. Shaw

Tars paper is a review of the principal details of cyanide practice in several of the modern plants in America, mainly during the year 1908. Two of the mills, the Goldfield Consolidated and the Vets, Colorado, are not yet prepared to give data, the former having started about Jan. 1, 1909, and the latter not yet being completed. These two mills embody most of the latest methods and are therefore described as far as possible. I. CRUSHING. The principal methods of crushing ores for treatment by the cyanide process are with gyratory and jaw-crushers, stamps, Bryan and Huntington mills, and tube-mills. The first stage is usually done by gyratory or jaw-crushers, in a few cases sup-\

Jul 1, 1909

By E. E. Donath

Application of the coal hydrogenation process for the production of chemicals is described. It has been estimated that a plant to produce 31,090 bbl per day of chemicals and fuels would cost $326,- 000,000. These products would consist of 6.1 pct by weight tar acids, 50.8 pct aroma tics, 16.4 pct LPG, and 26.7 pct gasoline. THE coal hydrogenation process is well known as a means for the production of liquid fuels from coal. In this paper the possibilities of the coal hydrogenation process as a source of chemical raw materials will be discussed. The recovery of chemicals from coal hydrogenation products necessitates some modifications and additions to the equipment used for the production of fuels. The conversion of coal into liquid fuels such as gasoline requires two steps: liquid phase and vapor phase hydrogenation. Intermediate oils obtained in the liquid phase contain oxygen and nitrogen compounds, especially aromatic compounds, since the coal itself is composed mainly of cyclic compounds. These oils are the raw material for the separation of high molecular weight aro-matics and of tar acids and tar bases. The relative amounts of heavy and middle oil—boiling above and below, e.g., 325°C—can be influenced by temperature, residence time, and catalyst of the liquid phase hydrogenation. In the vapor phase hydrogenation a solid bed catalyst is used to convert the intermediate oil from the liquid phase operation into hydrocarbons of the gasoline boiling range. Essentially pure hydrocarbons with a low content of sulphur, oxygen, nitrogen, and olefines are obtained. Depending mainly on the catalyst, the high ring compound content of the liquid phase oil can be preserved and gasoline rich in aromatics obtained. However, with other catalysts, these ring compounds can be split into gasoline with a high content of isoparaffins.' Again, depending on the operating conditions, products of varying boiling range can be obtained. Selection of catalyst and operating conditions therefore makes possible the production of different grades of motor fuel and also variations in the relative amounts of benzene, toluene, and xylene when aromatic hydrocarbons are produced. The schematic flowsheet, Fig. 1, shows the main steps of the coal hydrogenation process, and the product streams used for the recovery of chemicals are indicated. In the liquid phase hydrogenation stage, coal is made into a paste with recycle heavy oil. The coal paste is mixed with hydrogen and preheated at a pressure of 300 to 700 atm in heat exchangers and preheaters. The coal paste and hydrogen enter high pressure converters where the liquefaction of coal, conversion of the high boiling liquid materials initially obtained, into intermediate oils is performed at a temperature around 470°C. The reaction products enter the hot catchpot in which lower boiling oil vapors are separated from the residue containing unreacted coal, ash, and heavy oil. This residue, usually called H.O.L.D. (heavy oil let down), is processed to separate solid and liquid material. By thermal treatment of the residue, e.g., a coking process, distillate oil is obtained, which can be recycled as part of the pasting oil. This stream is rich in polynuclear aromatics and can be, and has been, used for the recovery of such material. The oil vapors and hydrogen from the hot catchpot are cooled in heat exchangers and a cooler, and the condensed oil is separated from the recycle hy-

Jan 1, 1953

By E. E. Donath

Application of the coal hydrogenation process for the production of chemicals is described. It has been estimated that a plant to produce 31,090 bbl per day of chemicals and fuels would cost $326,- 000,000. These products would consist of 6.1 pct by weight tar acids, 50.8 pct aroma tics, 16.4 pct LPG, and 26.7 pct gasoline. THE coal hydrogenation process is well known as a means for the production of liquid fuels from coal. In this paper the possibilities of the coal hydrogenation process as a source of chemical raw materials will be discussed. The recovery of chemicals from coal hydrogenation products necessitates some modifications and additions to the equipment used for the production of fuels. The conversion of coal into liquid fuels such as gasoline requires two steps: liquid phase and vapor phase hydrogenation. Intermediate oils obtained in the liquid phase contain oxygen and nitrogen compounds, especially aromatic compounds, since the coal itself is composed mainly of cyclic compounds. These oils are the raw material for the separation of high molecular weight aro-matics and of tar acids and tar bases. The relative amounts of heavy and middle oil—boiling above and below, e.g., 325°C—can be influenced by temperature, residence time, and catalyst of the liquid phase hydrogenation. In the vapor phase hydrogenation a solid bed catalyst is used to convert the intermediate oil from the liquid phase operation into hydrocarbons of the gasoline boiling range. Essentially pure hydrocarbons with a low content of sulphur, oxygen, nitrogen, and olefines are obtained. Depending mainly on the catalyst, the high ring compound content of the liquid phase oil can be preserved and gasoline rich in aromatics obtained. However, with other catalysts, these ring compounds can be split into gasoline with a high content of isoparaffins.' Again, depending on the operating conditions, products of varying boiling range can be obtained. Selection of catalyst and operating conditions therefore makes possible the production of different grades of motor fuel and also variations in the relative amounts of benzene, toluene, and xylene when aromatic hydrocarbons are produced. The schematic flowsheet, Fig. 1, shows the main steps of the coal hydrogenation process, and the product streams used for the recovery of chemicals are indicated. In the liquid phase hydrogenation stage, coal is made into a paste with recycle heavy oil. The coal paste is mixed with hydrogen and preheated at a pressure of 300 to 700 atm in heat exchangers and preheaters. The coal paste and hydrogen enter high pressure converters where the liquefaction of coal, conversion of the high boiling liquid materials initially obtained, into intermediate oils is performed at a temperature around 470°C. The reaction products enter the hot catchpot in which lower boiling oil vapors are separated from the residue containing unreacted coal, ash, and heavy oil. This residue, usually called H.O.L.D. (heavy oil let down), is processed to separate solid and liquid material. By thermal treatment of the residue, e.g., a coking process, distillate oil is obtained, which can be recycled as part of the pasting oil. This stream is rich in polynuclear aromatics and can be, and has been, used for the recovery of such material. The oil vapors and hydrogen from the hot catchpot are cooled in heat exchangers and a cooler, and the condensed oil is separated from the recycle hy-

Jan 1, 1953

By LIONEL E. BOOTH

ORE tests are made for the purpose of determining the correct methods of treatment for any particular ore. They should be conducted so as to insure that the results obtained in actual mill practice, on ore of similar grade and physical nature, will duplicate the results obtained in the test work. As the expenditure of many thousand of dollars may be involved in the building of the necessary mill, the testing of ores calls for extreme care and skill in handling laboratory machines, and a knowledge of actual practice to guard against the danger of obtaining results in test work that would be impractical to duplicate on a commercial scale. The importance of a correct sample of the ore is at once apparent. No matter with what degree of care the test 4rk is done, it loses its value when made on an unrepresentative sample. It is the first safeguard against future disappointments in the loss of capital and time. A number of failures in the past have been traced directly to faulty or lax methods in obtaining the ore sample.

Jan 1, 1924

By A. L. Sweetser

EXPERIENCE has shown how difficult it is to obtain information regarding laboratory-tests in connection with the chlorination-process for the extraction of gold from its ores, and I therefore present the following method, somewhat in detail, for the benefit of those who may desire to pursue research work in this field. The ore chosen was a partly-decomposed porphyry, extremely siliceous and comparatively soft, and of an average value per ton of from $25 to $26 in gold and $0.68 in silver. The chemical analysis gave : Fe2O3, 3.35; FeS, 1.5; MnO2, 0.75; ZnCO3, 4; Al2O3, 3.20; H2O, 0.7; SiO2 and insol., 83.50 per cent. The ore, first crushed to pass a 5-mesh screen, was coned and quartered until a 45-lb. sample was obtained. Samples for assay and analysis, taken from this lot by means of riffles, were crushed to pass a 100-mesh screen. The original sample was divided into 10-lb. lots, of which one was crushed to pass a 10-mesh screen; another through a 12-; the third through a 20-, .and the last through a 30-mesh screen. Because of the production of an excessive quantity of fine material it was deemed inexpedient to crush finer than 30-mesh size-a decision the wisdom of which is proved by the data in Table I., showing the increasing ratio of slimes, in proportion as the ore is crushed finer. In order to determine the percentage of fines in the sizes of mesh selected, a weighed sample of ore was taken from each mesh, and a sample from the 10-mesh lot was passed successively through 20-, 40- and 100-mesh screens, and the percentage remaining on each screen determined. This process was repeated with the 12-, 20- and 30-mesh lots, with the results given in Table I.

Sep 1, 1907

By Louis Raymond, John E. Dorn

The object of this investigation was to analyze the recovery that arises when the stress on a specimen undertaking creep is reduced. For this purpose annealed specimens of high-purity aluminum were precrept under a stress of 1000 bsi to a strain of 0.08 following which the stress was reduced for various periods of time to 10, 250, 500, or 700 psi. When the original stress was reapplied the subsequent creep curve lay above that for the unre-covered state and below that for the original annealed state. Analyses on the kinetics of this recovery as a function of the temperature gave a stress-sensitive activation energy that decreased as the reduced stress was increased from a value of 64,000 cal per mole at 10 psi to 37,000 cal per mole at 750 psi. Recovery was also detected and measured during creep under the reduced stress. Following a short initial period, the creep rate under the reduced stress increased monotonically until it reached the secondary-creep rate for the reduced stress. The temperature dependence of this phenomenon was also shown to be correlatable in terms of the previously deduced activation energy for recovery. The activation energies for creep of most pure metals at high temperatures have been shown to agree well with those for self-diffusion.'j2 Since the true secondary stage of creep is usually due to the steady-state balance between the rate of strain hardening and the rate of recovery, it is generally thought that the activation energy for recovery of the creep-induced substructure equals that for creep itself. A shoft time ago, however, Ludemann, Shepard, and Dorn~ found that the activation energy for recovery of the creep-induced substructure in high-purity aluminum under zero stress was almost twice that for self-diffusion, namely about 65,000 cal per mole; obviously recovery under reduced stresses differs in some significant way from the recovery that accompanies the secondary stage of creep. The major purpose of this investigation is to study the effect of stress on the re- covery of the creep-induced substructure in order to provide a better understanding of the recovery mechanism itself. EXPERIMENTAL TECHNIQUE High purity aluminum, containing 0.004 pct Cu, 0.002 pct Fe, and 0.001 pct Si, used in this investigation, was in the form of 0.100-in.-thick sheet which has been cold-rolled to the H-18 temper. Creep specimens were milled from the sheet with their tensile axes in the rolling direction. All specimens were then heated at 686°K for 1 hr followed by air cooling in order to produce an annealed structure which exhibited a uniform equiaxed grain size of about 4 grains per mm. Tests were run in creep machines fitted with Andrade-Chalmers type of lever arms so contoured as to maintain the stress constant to within 0.05 pct of the reported values. Constant temperatures to *O.l°K were obtained by complete immersion of each specimen in a temperature-controlled and agitated bath of molten KN02-KNOs mixture. Where changes in temperature were involved, the change was effected in less than 2 min by manually replacing one bath by another controlled at the second temperature. Displacements over the gage section were sensed by linear differential transformers, the output of which was autographically recorded. The calculated strain measurements were sensitive to 5x EXPERIMENTAL PROCEDURE The following analyses are based on extensions of the previously announced effect of the temperature on the creep strain,2 namely for a = constant, where e = the total true tensile creep strain for a given applied true tensile stress, t = the duration of the test, R = the gas constant, T = the absolute temperature, Q, = the activation energy per mole for creep which is independent of the stress, / = a function of 8, = and of the stress, and a = the stress. The validity of this correlation for high-purity aluminum is demonstrated in Fig. 1 for temperatures in the near vicinity of 600°K; the activation energy for creep, Q,, which is approximately that for self-diffusion, is insensitive to the applied stress

Jan 1, 1964

By Donald A. Dahlstrom

In 1961 it became increasingly more evident that significant technological changes must be made in minerals beneficiation. The status quo in technology in today's competitive markets will not be tolerated. Higher productivity, greater recoveries and newer and better products must be sought and incorporated with lower operating costs. In other words, "when you build a new mill, don't build an old mill." Recommended reading for proof of the vital necessity of new methods and techniques is in the September 1961 issue of MINING ENGINEERING. The article by one of MBD's cherished and true veterans, E. H. Rose, is entitled "Iron Ore: The BIG Picture". Because of new beneficiation methods for iron ores, nearly 47 million tons of plus 60% iron concentrates will be produced and another 40.1 million tons of capacity will be under construction or committed. Thus in a relatively short period, 87 million tons of this high-grade iron concentrate with record low silica will be available. As Mr. Rose also points out, "in gross tons, without allowance for better grade, this is the equivalent of almost exactly two-thirds of all iron ore used for all purposes in the U.S. and Canada in the peak years of 1953 or 1957." Needless to say, improved mineral beneficiation methods have had far more reaching effects than just the production of higher grade iron concentrate.

Jan 2, 1962

By R. A. Wolfe, H. W. Paxton

Self-diffilsion coefficients for cr51 and Fe55 in 12 pct Cr-Fe and 17 pct Cr-Fe for Fe55 in chromium, and for Cr51 in vanadium have been measured. The results are compared with other values for the Fe-Cr system, and with the various theories of diffusion in hcc metals. Some empirical correlations are discussed between Do and Q in hcc systems, or, expressed differently, the constancy of ?G*/T solidus for seveval bcc metals and alloys is noted. It appears very probable that a vacancy mechanism is operative in bcc metals, hut this cannot he stated with certainty. THE great bulk of work on diffusion in metals, both experimental and theoretical, was for many years concentrated on those with close-packed and, in particular, fcc lattices.1,2 There appears to be little doubt that the mechanism of diffusion in these solids is vacancy migration, leading to mass transfer and in substitutional solid solutions to a Kirken-dall effect.3,4 For bcc metals, the picture is much less clear. The Kirkendall effect certainly occurs in several alloys.5-10 However, attempts to understand the factors contributing to the pre-exponential in the usual expression for the diffusion coefficient D =D, exp {-Q/RT) by extension of ideas useful in close-packed lattices have not always been successful. Zener,11 Leclaire,12 and Pound, Paxton, and Bitlerl3 have suggested that various forms of ring diffusion may be important in some bcc metals. For close-packed metals, Do is usually about 1 sq cm per sec and Q - 35Tm kcal per mole (Tm = melting temperature in OK). The theory of Pound et al. suggests for ring diffusion that Do may be about 10-4 and Q, although difficult to calculate with any precision, would be significantly less than 35 T,. The experimental results on self and solute diffusion in ? uranium14,15 and ß zirconium,10 and for solutes in 0 titanium,17 and possibly for self-diffu- sion in chromium below about 0.75 T,," gave some credence to this theory. However, not all bcc materials display low values of DO and Q, and the exceptions were not predicted by any theory. Furthermore, it has recently become apparent that, in bcc materials, log D is not always linear with T-l if a sufficiently wide range of temperature is studied.16,18 This variation may be such that Q may increase18,19 or decrease20 with increasing temperature. The present work was undertaken in an attempt to provide further diffusion data on bcc metals, and to try to understand the factors which contribute to differences in behavior between the various elements. For part of this work, the Fe-Cr system was chosen since it is of considerable technological importance, and data on 12 pct Cr and 17 pct Cr alloys appeared well worthwhile to supplement that existing for the remainder of the stern.18,22 The diffusion of Fe55 in chromium was studied as an example of a more or less "normal" tracer element in a possibly abnormal host lattice. Finally, no data were available for vanadium, the neighbor of chromium in the periodic table, because of lack of a suitable isotope so cr55 was used as a tracer in a few preliminary experiments. For convenience, we shall refer to elements whose Do and Q are low compared to those predicted by Zener's theory as "anomalous". PROCEDURE This investigation determined self-diffusion rates by means of radioactive tracers and the integral-activity method first utilized by Gruzin.23 In this method a thin layer of radioisotope of the diffusing element is plated or coated onto a planar surface of the diffusion sample, which is then given an isothermal-diffusion annealing treatment. The determination of an activity-penetration curve involves measuring the residual activity of the specimen after each successive layer or section has been removed parallel to the original planar surface. The method used here is essentially the same as that used by Gondolf18 and Kunitake.21 Two radioactive tracers, cr51 and Fe55, were used in this investigation. Diffusion coefficients were determined for the diffusion of one or both of these tracers in four different materials, viz., Fe-12 wt pct Cr alloy, Fe-17 wt pct Cr alloy, chromium, and vanadium. The diffusion samples had nominal dimensions of 1.5 cm diameter and 0.5 cm thickness. The grain size was several millimeters for the Fe-Cr alloys and at least 1 mm for the chromium and vanadium samples. Accurately planar surfaces

Jan 1, 1964

By Edward Griffith

The anthracite industry, which produces about 50 million net tons of coal annually, has been talked of as being able to last for another century; but if the water record of the past century continues into the expected century, this expectation of many years production still to come out of the remaining reserves will not be attained. The industry will be a submerged one in fact as well as in name. A recorded forecast of what has since occurred appears on page 34 of the Mine Inspectors' reports of 1897: It should also be noticed that bodies of water had accumulated in parts of abandoned mines before duplicate surveys of the same were required by law; and as a result we have today to contend with bodies of water, the exact location and position of which are not correctly known. Since that date, 47 years ago, the implications of the foregoing comment have grown enormously. This is shown in Fig. I and in the following table showing flooded and partially flooded areas that have come into existence since the early 1930's in the northern anthracite field: Number of Year Flooded Areas 1932...................... 2 1933...................... 3 1934..................... 3 1935...................... 5 1936...................... 7 1937...................... 9 1938.................... 11 1939...................... 11 1940...................... 11 1941...................... 14 1942...................... 14 1944...................... 26 Similar conditions exist in other fields in the anthracite region. In the period between 1923 and 1940, abandonment of collieries in the southern field resulted in the flooding of approximately 60 sq. miles (30 per cent) of the field. The ahandonment of collieries in this field has imposed additional expense for pumping on the collieries still operating. This condition will be aggravated by the extension of developments and cause suspension of operations if economic conditions do not justify the assumption of additional pumping costs. During the years, the region's rainfall has not altered, but the quantity going inside the mines through breakages in the surface, brought about by extraction of the coal, has increased materially, as indicated by the ratio of tons of water pumped to tons of coal produced underground for one of the large companies in the northern anthracite field: Year Ratio 1920................. 8.4:1 1925.................. 10.6:I I930.................. II.4:I 1935................. 26.2:1 1940.................. 32.7: I 1942.................. 30.3:I The data in Table I shows the total amount of water pumped and its ratio to underground production for one of the large companies in the southern anthracite field, and shows a rapidly growing serious situation. Rainfall quickly affects the workings in all the fields. In mine inspection district 7 of the southern field the rainfall in 1935 was 92,083,000,000 gal. The total water pumped was 36,274,000,000 gal. (39.4 per cent of the rainfall). The operating mines

Jan 1, 1947

By Edward Griffith

The anthracite industry, which produces about 50 million net tons of coal annually, has been talked of as being able to last for another century; but if the water record of the past century continues into the expected century, this expectation of many years production still to come out of the remaining reserves will not be attained. The industry will be a submerged one in fact as well as in name. A recorded forecast of what has since occurred appears on page 34 of the Mine Inspectors' reports of 1897: It should also be noticed that bodies of water had accumulated in parts of abandoned mines before duplicate surveys of the same were required by law; and as a result we have today to contend with bodies of water, the exact location and position of which are not correctly known. Since that date, 47 years ago, the implications of the foregoing comment have grown enormously. This is shown in Fig. I and in the following table showing flooded and partially flooded areas that have come into existence since the early 1930's in the northern anthracite field: Number of Year Flooded Areas 1932...................... 2 1933...................... 3 1934..................... 3 1935...................... 5 1936...................... 7 1937...................... 9 1938.................... 11 1939...................... 11 1940...................... 11 1941...................... 14 1942...................... 14 1944...................... 26 Similar conditions exist in other fields in the anthracite region. In the period between 1923 and 1940, abandonment of collieries in the southern field resulted in the flooding of approximately 60 sq. miles (30 per cent) of the field. The ahandonment of collieries in this field has imposed additional expense for pumping on the collieries still operating. This condition will be aggravated by the extension of developments and cause suspension of operations if economic conditions do not justify the assumption of additional pumping costs. During the years, the region's rainfall has not altered, but the quantity going inside the mines through breakages in the surface, brought about by extraction of the coal, has increased materially, as indicated by the ratio of tons of water pumped to tons of coal produced underground for one of the large companies in the northern anthracite field: Year Ratio 1920................. 8.4:1 1925.................. 10.6:I I930.................. II.4:I 1935................. 26.2:1 1940.................. 32.7: I 1942.................. 30.3:I The data in Table I shows the total amount of water pumped and its ratio to underground production for one of the large companies in the southern anthracite field, and shows a rapidly growing serious situation. Rainfall quickly affects the workings in all the fields. In mine inspection district 7 of the southern field the rainfall in 1935 was 92,083,000,000 gal. The total water pumped was 36,274,000,000 gal. (39.4 per cent of the rainfall). The operating mines

Jan 1, 1947

By Kenneth L. Cook

GEOPHYSICAL exploration is continuing to expand in scope and variety of methods, in experimental studies of specific applications, and in development of theory for complex situations. Aerial surveys have been made economically more attractive by simultaneous recording of magnetic and radioactivity data, and aerial electromagnetic surveys have been attempted. The most significant success of mining geophysics was the discovery of huge lead-zinc-copper sulphide ore deposits in New Brunswick, Canada. Spectacular geophysical astivity has occurred in the area as a result of these discoveries, and geophysics is expected to continue to play a large role in this drift-covered area. Other developments this year find the Atomic Energy Commission making available monthly reports of the location of anomalous radioactivity detected in aerial and ground surveys and the AEC is installing supersensitive radioactivity assaying instruments in about 25 localities in the U. S. Interest in helicopters is being aroused because of potentialities for near-ground surveys and better location control of anomalies. Japan can be added to the list of countries using geophysical surveys extensively.

Jan 2, 1954

By J. Chipman, G. G. Hatch

One of the important functions of the iron blast furnace is the desulphur-ization of pig iron before it enters the steelmaking furnaces. However, the increasing concentrations of sulphur in the metallurgical coke, source of approximately 90 pct of the sulphur present in the blast furnace charge, and demands for higher rates of production within recent years have increased the need for greater desulphurization within the iron blast furnace. Furnace operators are beginning to look for desulphurizing agents other than blast furnace slag to accomplish the desired degree of desulphurization. A considerable amount of work has been done on desulphurization outside the furnace with soda ash, calcium carbide and various synthetic slags. Whether the desulphurization of pig iron is accomplished wholly inside the furnace or partly inside and the remainder outside, will be determined by the economics involved. Regardless of which is the case, it is believed that it is necessary to have a better understanding of the physical chemistry of desulphurization by blast furnace slags. To this end, it is the object of the present investigation to attempt what is believed to be the first equilibrium study of the distribution of sulphur between liquid pig iron and a wide range of blast furnace slag compositions. Review of Literature There is a considerable amount of information in the literature concerning the desulphurizing power of iron blast furnace slags, the solubility of various sulphides in the slags, and the effect on desulphurization of temperature, of elements dissolved in the liquid iron, and of viscosity. However, there is nothing to indicate that the equilibrium distribution of sulphur between liquid iron saturated with carbon and iron blast furnace slags has been studied experimentally. Wentrupl has made probably the most detailed study of the desulphurization of pig iron to date. He considered that there are three distinct aspects involved, namely: 1. Desulphurization within the blast furnace (by lime and manganese). 2. Subsequent desulphurization by manganese. 3. The effect of subsidiary reactions on the desulphurization by manganese. The experimental work carried out by Wentrup was devoted mainly to obtaining a better understanding of how desulphurization by manganese was accomplished in the mixer and the ladle. Particular attention was given to the part played by carbon, silicon, and phosphorus associated with manganese in the iron, and the effect of temperature on desulphurization. The experimental results indicated that desulphurization by manganese is purely a process of crystallization of manganese sulphide. The addition of silicon to iron melts containing 3.5 pct carbon and less than 0.5 pct manganese had no noticeable effect on desulphurization, but with 1-2 pct manganese the silicon additions improved the desulphurization. Additions of phosphorus also resulted in improved desulphurizati011 by manganese, but the effect was not as marked as in the case of silicon. It was also found that desulphurization by manganese was further improved by lowering the temperature. In order to explain desulphurization inside the blast furnace, Wentrup considered the system iron, sulphur, calcium, oxygen, manganese. (silicon). The distribution of sulphur between the metal and slag was represented by the following equation: (SS) _ (S)Fe + (S)Ca + (S)Mn .... [S] = [s] [1] The parentheses and the brackets represent the equilibrium concentrations in weight per cent of the slag and metal constituents, respectively. Since FeS D (FeS) _ (FeS) (S)Fe LfeS - [FeS] [S] [2] (CaO) + S e (FeO) + (S)Ca _ (FeO)(S)Ca. (S)Ca _ (CaO) Kl = (CaO)[S] [S] ~K1(FeO) [3] Mn + S D (S)Mn (S)Mn (S)Mn K' = [MnpT "1ST = *lIMnJ !4) Substitution of Eq 2, 3, and 4 into Eq 1 resulted in if = L- + *> (Sol + K^ (S) [51 Eq 5 was used to calcu1;lte -f^j and [S] at 1480°C for slags containing 30-50 pct lime, 0.1-2.5 pct iron oxide, 0-26 pct silica, 2 pct sulphur and iron analyzing 1.5 pct manganese. The value for LFaB at 1480°C was found to be equal to 4.5, based on the experimental work of Bardenheuer and Geller.2 The results of the calculations are shown in Table 1. Although the slags are hypothetical and do not represent the range of compositions found in ordinary blast furnace practice, the calculations indicate that lime is effective in controlling desulphurization only if the iron oxide and silica contents of the slag are kept low. Schenck3 did not claim K1 to be a true equilibrium constant, but an empirical value which varied with the silica content of the slag.

Jan 1, 1950

By A. P. Watts

FOREWORD THE problem of concentrating the disseminated lead ore of southeast Missouri is extremely simple. The economic mineral is galena and the gangue is dolomite. The ore assays from 4 to .6 per cent. lead, 80 per cent. of which is recovered in a 70 per cent. concentrate.. The ore is crushed through 10 mm. and sized on 2 mm.; the oversize is jigged, while the undersize is tabled, and the slime treated by flotation. The district is the largest lead producer in the world, the output of metallic lead for the year 1915 being 183,906 tons. Over 20,000 tons of ore a day is being treated. The management of the different properties is progressive and always open to suggestion. The fact; however, must not be lost sight of, that the problem in the district is peculiar, as the average ore contains but 80 to 120 lb. of a metal normally selling at about 4c. per lb. The low price of the metal limits the treatment system to one of low cost. An elaborate treatment is neither logical nor advisable. An engineer first visiting the district may consider that the method of treatment could be improved

Jan 10, 1917

By H. J. Ramey, I. S. Bousaid

Experimental results on the oxidation reaction kinetics in the forward combustion oil recovery process are presented. A total of 48 runs were made wherein a stationary thin layer of coked, unconsoli-dated sand was burned isothermally in a combustion cell. Individual runs were made at various temperature levels to permit determination of the effect of temperature upon the reaction. An expression was obtained for the burning rate of carbon as a function of carbon concentration, combustion temperature and oxygen partial pressure. The carbon burning rate for two types of crude oil indicated a first order reaction with respect to both carbon concentration and oxygen partial pressure. The effect of combustion temperature on the reaction rate constant matched the Arrhenius equation. The activation energy was similar for the two crude oils examined. The activation energy decreased for a porous media containing clay. The rate of oxidation of crude oil at reservoir temperature was found to be significant. Other significant findings included information on hydrogen-carbon content of fuel residues, fuel reactivity and the products of combustion. INTRODUCTION The production of crude oil by underground combustion has been studied in the laboratory by many investigators.1-5 Results of laboratory and field experiments have been reported in the literature describing the forward combustion process. But as yet, no qualitative or quantitative study of the kinetics of fuel combustion involved in this process has been reported. The fuel concentration and the rate at which fuel is burned at the front are important factors governing the air requirement in a forward combustion operation. Although the fuel is essentially unrecoverable crude, the air required to burn the fuel is an important economic factor in this process. Because fuel is burned, the heat transport associated with forward combustion is a key and unique feature of this new oil recovery method. Many investiga-tors5-9 have presented information on the heat transmission and fluid mechanics involved in forward combustion. Berry and Parrish10 demonstrated the utility of considering reaction kinetics in reverse burning. From differential thermal analysis, Tadema11 presented a qualitative discussion of the nature of reactions between oil and oxygen in combustion oil recovery. Although little quantitative work has been done on the reaction kinetics involved in forward combustion oil recovery, an extensive literature does exist on combustion of carbons and oils, and carbonaceous residues from cracking catalyst pellets. Dart, et al., 12 studied the combustion rate for oxidation of carbonaceous residues on clay catalyst pellets, and found the reaction to be second-order with respect to carbon concentration, and first-order with respect to oxygen partial pressure for carbon concentrations less than 2 weight percent of the catalyst weight. The reaction appeared to be first-order with respect to carbon concentration for concentrations greater than 2 percent. Metcalfe13 noted that other workers had found that aging of the fuel during the combustion process was responsible for changing coke properties, and it accounted for the apparent second-order carbon concentration effect found by Dart, et al. It appears that burning of residues from cracking pellets is first-order with respect to both carbon concentration and oxygen partial pressure. Dart, et al., also observed that hydrogen in the hydrocarbon residue appeared to react faster than the carbon. Lewis, et al.,14 studied oxidation of charcoal, coke and graphite in a fluidized bed. Gas velocities were high enough to partially lift and circulate the carbon particles. Their results indicated first-order reaction dependency with respect to both carbon concentration and oxygen partial pressure. One

Jan 1, 1969

By Baxter C. Madden

Stress corrosion is defined and certain examples are illustrated. A number of test methods, to determine the resistance of materials to stress corrosion, are being developed at Wright Field. The criterion as to the validity of a method is stated. Three types of specimens and the method of stressing each type are described. Five different corrosive conditions are set forth. The results of tests to date, although too few for any general conclusions to be arrived at, are presented. The program for the continuation of the investigation is indicated. A bibliography is included. The development of the method of loading a cantilever-type specimen, so that a nearly constant stress at a large deflection may be realized, is presented as Appendix I.

Jan 1, 1945

By Alistair Frame

INTRODUCTION In April 1988 discussions were about to start with BP on the possibility of buying their- mineral interests, and it took a long time to arrange a deal. I will talk about RTZ without BP Minerals, look at the new RTZ with BP Minerals, discuss the outlook for metal prices, and review the problems in South Africa and develop the case against sanctions on that country. RTZAT THE END OF 1988 Late in 1987, after completing a worldwide review of our strategic thrust, we decided: 1. That the business focus would be to remain a major diversified natural resource group with interests in related industries. 2. To divest non-core and low growth businesses. 3. To develop a logical acquisition plan. 4. To emphasise organic growth and, most importantly, to enhance shareholders' wealth by capital increase and a progressive dividend policy. Subsequently, we concluded that we should sell our oil and gas business and dispose of several assets for which we saw little growth potential. The disposal programme raised some U.S.$ 1,600 million, representing a significant capital gain on our various investments. All transactions were carried out by our small Head Office staff in London, without the help of merchant banks and lawyers, and we were able to safeguard the employment of our staff in all these operations. The most interesting disposal was of our oil and gas interests. Before the 1987 strategic review, we had ambitions to develop our oil and gas interests and to become the leader in the second division of European oil companies. Our interests, which were purely in exploration and production of oil and gas, consisted of the 100% owned RTZ Oil and Gas company, and 29% of LASMO, a publicly quoted U.K. oil and gas company with extensive worldwide interests. LASMO, in turn, owned 25% of Enterprise Oil, one of the more successful of the North Sea oil exploration and production companies. Prior to the end of 1987, our intention was to acquire control of both LASMO and Enterprise. However, when we evaluated the net asset value of these two companies based on our oil price assumptions, which were pessimistic for the next five years or so, we concluded that they were both trading in the market at significant premiums. This, taken with the additional premium required for acquisition and the high cost of exploration, led us to conclude that, since funds could undoubtedly be used to better effect, we should sell our interests. Turning to the other major achievements in 1988, we commissioned, and built up to design levels in a very short period of time, the Morro do Ouro gold mine in Brazil. This mine reached its design output of 100,000 ounces of gold per year within three months of start-up, which was on time and under budget. We are now considering a possible expansion of this property. In South America, Utah International, Mitsubishi and RTZ arranged non-recourse finance, with post completion guarantees, for the giant Escondida Project in Northern Chile. Financing came from a group of banks in Japan, Germany and Finland -- the countries to which most of the concentrates will be sold -- and from the LEC. The average output of the mine for the

Jan 1, 1990

By J. W. H. Clare

An equilibrium isotherm at 550°C is given for ternary alloys rich in aluminum containing 0 to 15 wt pct Cr and 0 to 20 wt pct Mn. Phases encountered are: aluminum solid solution; stable temary compound G; and MnAl8,. In addition, a limited number of alloys were annealed at 450°C to check for differences from the 550°C isotherm. These experiments limited the composition of the G phase at 550°C and further experiments established that the G Dhase is not stable above 590 ± 1 °C. RAYNOR and Little1 established several equilibrium isotherms using aluminum-rich alloys containing up to 5 pct of the elements chromium and manganese. These isotherms showed that, in addition to an a solid solution and the two binary intermediate compounds 8 and MnAl6, a stable ternary compound ex-isted in the system. This phase was called the G phase because of its ghostly appearance in micro-sections. Little, Raynor, and Hume-Rothery2 found a similar phase in binary aluminum-manganese alloys and, as a result of their experiments, concluded that the G phase was a metastable phase existing in the aluminum-manganese system. This phase was stabilized by the addition of chromium. The constitutional work of Raynor and Little,' although showing the existence of the G phase in the aluminum-chromium-manganese system, did not establish precisely the composition limits of the phase or the temperature above which the phase became unstable and disappeared from the system. The work described in this paper was undertaken to provide this information. An examination of the previous work' showed that an isotherm at 550°C would provide the required data regarding composition in the shortest possible time, as the (a + G) phase field narrows considerably at higher temperatures. The same work showed that the G phase ceased to exist between 580° and 595°C. A more precise estimate of the transformation temperature would therefore be obtained by annealing selected alloys at successively higher temperatures beginning at 580°C. EXPERIMENTAL PROCEDURE 1) Materials Used—The experimental alloys were prepared from four master alloys containing 15.00 w pct Cr, 15.28 wt pct Cr, 19.80 wt pct Mn, and 19.38 wt pct Mn which were made from super-purity aluminum, electrolytic chromium, and electrolytic manganese. The aluminum was of 99.996 pct purity, the principal impurities being copper, iron, and silicon. The chromium and manganese were 99.9 pct pure, the principal impurities being antimony, calcium, and magnesium. 2) Preparation of the Experimental Alloys—The alloys were prepared in 20-g quantities by melting together the requisite amounts of master alloy and aluminum in an alumina-lined crucible and, after thorough stirring with an alumina rod, cast into a cold copper mold to give ingots 1/4 in. in diam and 3 in. in length.

Jan 1, 1960

By Stanley Reichert

The manganese deposits of the Embree Iron Co., Embreeville, Tenn., are thought to be fairly typical and representative of the East Tennessee type of deposit, and so to warrent rather detailed description. The methods used by the Embree Iron Co. for prospecting and mining, and to a lesser extent for milling; should therefore be applicable to other East Tennessee manganese deposits, as yet' undeveloped. Early History and Discovery of Manganese Brown iron ore (limonite) was mined by the Embree Iron Co. at various times from 1890 to 1910. The early method was simply pick-and-shovel mining in open pits. Later hydraulic giants were used and sluice lines carried the ore down the mountain slopes to log washers.. During this phase of mining, manganese ore was discovered under the iron ore in some of the pits. In other areas oxidized zinc ore was found under the iron. It thus happened that the valuable zinc deposits were discovered in 1914. During the World War period, from 1914 to 1918, there was sporadic prospecting for manganese ore among the abandoned iron workings, but the main activity then centered around the recently discovered lead and zinc deposits. It was not until 1934 that manganese mining began in earnest, under the supervision of R. D. Pritchett, Mine Superintendent. Geology of the Manganese Deposits United States Geological Survey Folios 151 (Roan Mountain) and 118 (Greeneville show that Bumpass Cove is a synclinal structure thrust faulted along the southeast flank, but with normal sequence of strata along the northwest flank. Fig, l. shows the areal geology of Bumpass Cove as taken from the folios superposed on the U.S.G.S. -T.V.A. 1/24,000 topographic maps of the Erwin and Telford 7 ½ minute sheets. The manganese mines occur near the contact of the unfaulted Shady dolomite and the Erwin quartzite. This contact zone is deeply weathered, with overburden averaging 40 to 60 ft. thick. The material overlying' the bedrock along this normal contact is a very tough, plastic clay, termed "buckfat" or "ocher" by the miners. It is not a true ocher, however, in the sense of having commercial value. The manganese deposits occur along this "ocher" band. Fig. 1 also shows how the manganese deposits occur along the broad topographic terrace between the 2100 and 2200-ft. contour lines. This terrace lies at the foot the Erwin quartz ridge. Apparently it is the topographic is expression of the deeply weathered zone along the Shady-Erwin contact.

Jan 1, 1941

By Eckley B. Coxe

I wish to call the attention of the Institute to a peculiar variety of anthracite which occurs in the Buck Mountain vein at our collieries at Drifton, and in the same and other veins in different localities in the anthracite coal-fields. It is known among the workmen as iron-gray or cast-iron. It has a dull, greasy appearance, as if a piece of ordinary anthracite had been rubbed with a dirty and greasy rag and allowed to dry. It was formerly considered a very impure coal containing a high percentage of ash, and was picked out of the coal and thrown away. But the following analysis of it and of its ash, made by Mr. J. Blodget Britton, of Philadelphia, shows that such is not the case, the percentage of ash not being much higher than that of the ordinary anthracites as they are prepared for market. ANALYSIS OF COAL. Water,.........4.36 Fixed carbon, by difference,.. 84.90 Volatile combustible matter,. . 3 48 Ash,..........7.26 100.00 The carbon proved to some extent to be of the nature of graphite, and slowly combustible. ANALYSIS OF ASH. Silica,.........25.66 Sulphur,........ .17 Alumina,........27.03 Phosphoric acid,.......21 Ferric oxide,.......42.83 Alkali, undetermined matter and Lime,.........1.56 loss,.........60 Magnesia,........1.83 Manganous oxide,......11 100.00 We recently collected a quantity of the substance and made a fire with it, in an ordinary stove, which was kept up several days. We found that it burned with intense heat, and apparently as freely as our other coal does. Mr. Britton is probably right in his view, that the difference in appearance is not due to impurities in the coal, but to the fact that the carbon is in a more or less graphitic form, as many of the standard coals are no purer. I bring this before the Institute as I know that a large quantity of this valuable fuel is thrown away every year, on account of its supposed impurities.

Jan 1, 1879

By A. J. Boynton

THIS paper is not the result of recent research with regard to any particular feature of iron metallurgy, blast-furnace practice or mechanical engineering. It is rather a series of notes with regard to contemporaneous progress in blast-furnace construction and operation, together with some suggestions as to changes and improvements that seem to be desirable. On April 1, 1930, a committee of the Chicago District Blast Furnace and Coke Association presented to that association a rather ambitious study of blast-furnace equipment, together with recommendations for the design of construction of a complete blast-furnace unit for the production of 1000 tons of pig iron per day. This report represented the result of months of study by members of the committee, all of whom were actively engaged in operation or design of blast-furnace plants. It is natural, therefore, that comments on the present state of the art should take this 1000-tons furnace, as recommended by the Association, as a basis for comparison with corresponding features of construction or design. See Table 1. In looking through this report at the present time, the significant difference is in the nature of the gas-washing equipment that today would be specified. In other respects, the installation would remain generally very much as it was outlined in 1930. The change in furnace dimensions would be very slight and of no especial significance, except with respect to changes in stock-line diameter, of which more will be said hereafter. The high-efficiency stove as developed by various designers is thoroughly described within this report, and few changes have taken place since. Blowing equipment shows refinement and reflects the improvement in steam engineering. Hoisting and charging equipment and ladles for the conveyance of iron and slag remain, in general, as they were in 1930. In accordance with a general tendency

Jan 1, 1935