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By R. E. Hudrlik, G. W. Preckshot

The diffusion coefficients of silver from solid silver in molten lead were measured to within ± 0.8 pet in a columnar type diffusion cell ower, the temperature range of 326° to 530°C. Fick's law describes the process up to 530°C where the laminar mechanism appareltly breaks down. These is negligible resistance at the interface as shown by mathematical analyses. The diffusion coefficients are found concentration independent. IT would seem that diffusion in liquid metals would be free of such effects as molecular structure, dissociation. polarization. and compound formation. This view was taken by Gorman and preckshot in their study of diffusion of copper from solid copper into molten lead. They reported diffusion coefficients which were independent of the concentration over the range of 478° to 750°C. They found that the Stokes-Einstein equation with constant radius of the diffusing specie represented the diffusion data better than Eyring's rate theory equation and Sheibel's correlation. The radius of diffusion was found to be that of the doubly charged copper. There appeared to be no resistance across the solid-liquid boundary. In the present work the diffusion coefficients for silver in liquid lead were measured over a range of temperatures of 350° to 505°C. The solubility of silver in lead over the range of 303° to 630°C was also obtained. These results are compared with calculated or correlated values or with data in the literature. EXPERIMENTAL Procedure—The experimental equipment techniques and procedures were those reported in detail by Gorman and preckshot9 and will not be repeated here. Measured values of WT, Co, A. L were obtained for various diffusion times and the diffusion coefficient was computed for the case of no resistance at the interface9, 11 by: WT/CoAL = 1- 8/p2 n=1 1/(2n - 1)2 exp[-(2n - 1)2p2 Dt/4L2] [1] or where there was resistance at the interface by: WT = 1- ?n=1 2h2/ap2L [sxp [-Dan2t]/[(h2 + an2) L + h] The roots an are those of the transcendental equation3 tan (an L) = Iz/cun. The diffusion coefficient is that defined by Hartley and Crank.7 The total silver in the lead cylinder and equilibrium slug was determined by a cupellation technique' with proper correction for losses. Analysis of known samples showed that this method is surprisingly accurate. The amount of silver in the lead adhering to the silver cylinder was obtained in the same fashion as shown by Gorman and preckshot.9 The small errors involved in this determination are not critical since the silver in this adhering lead layer is only 3 to 15 pet of the total diffused. Materials—Electrolytic silver containing 99.9+ pet Ag obtained from General Refineries of Minneapolis, Minn. was used for all but runs 7 and 8. For the balance of the runs this silver was reduced with hydrogen at 1100°C and its oxygen content was found to be about 0.017 pet. For the runs. 7 and 8, phosphorous-reduced silver of the same purity was obtained from Handy and Harman Co. of Chicago, Ill. The densities of the phosphorus-reduced silver and the hydrogen-reduced electrolytic silver were 10.484 and 10.487 g per cm3, respectively. These values agree with those reported for pure silver. Silver which was reduced at 900 C had an average density of 9.998 g per cm3, indicating porosity. This silver was used for a number of runs which were not tabulated in Table I. These are indicated by crosses on Fig. 2. The 99.999 pet Pb was obtained from the National Lead Co. Research Laboratory of Brooklyn, New York. DISCUSSION OF RESULTS The diffusion and solubility results are reported in Table I for eleven runs using either phosphorus-reduced electrolytic silver or hydrogen-reduced silver at 1100° C. The solubility data shown in Fig. 1 show the excellent agreement with that reported by Heycock and Neville.8 The data of Friedrichs5 apparently are in error. The experimental solubility data of this work are reported to 0.3 pet. The experimental diffusion coefficients computed from Eq. [1] are reported within 1.2 pet of the mean and are plotted in Fig. 2. These are expressed within +0.8 pet of the experimental values over the entire temperature range by: D= 8.26 x 10 -5 e-1925/RT . [3] There appears to be little difference due to the

Jan 1, 1961

By T. M. Kegley, J. H. Frye, D. L. McElroy, M. L. Picklesimer, E. E. Stansbury

Measurements of rate of growth, thermodymmic quantities, and partitioning of Mn are reported for high-purity eutectoid Fe-C and Fe-C-Mn steels for the auistenite-pearlite reaction. Evaluztion of the cotnstants of a derived-rate equation with these data indicate that Mn additions cause the decrease observed by increasivg the activation energy and apparent activation entropy or the movenment of complexes at the interface. A diffusion-controlled mechanism is not required. Appreciable partitioning of Mn to cementite occurs entirely behind the austenite-pearlite interface for more than 30°C of subcooling. ThE growth of pearlite from austenite has been extensively studied over a period of years. but the rate-controlling step in the process is not known nor is the mechanisrn of the effect of alloying elements understood. A comprehensive review of the subject has been given by Mehl and Hagel. The present study has been concerned with further considerations of the effects of alloying elements on the rate of growth of pearlite. Analytical relatioriships for the rate of growth of pearlite have been considered by Zener.2 Turnbull,3 Machlin,4 Cahn.5 Frye,6 and Frye et a1.7 The derived expressions in each case involve the free energy of the transformation. the interlamellar spacing. and some rate term of the nature of the diffusion coefficient. mobility or frequency of occurrence of some rate-determining step. The treatments. of course. differ in detail. Further consideration is given in the present paper to the equation of Frye, Stansbury. and McElroy which was derived from modified absolute rate theory without in any way specifying the mechanism of the reaction. This equation is r=c ? T ? F exp -?E*/RT [1] where r = rate of growth in mm/sec c = a constant AT = equilibrium transformation temperature minus actual transformation temperature "C ?F = free energy per mol of austenite minus that of pearlite ?E* = activation energy R= gas constant ?F in Eq. [1] is computed from the following: ?F=?H?T/To - ?T' dT' ?T' ?CdT"/T" [2] To To where AH is the enthalpy per mol of austenite minus that of pearlite at the equilibrium temperature, To, and AC is the difference in specific heat between austenite and pearlite per mol. Subsequently, Darken, Fisher, and Galligan8 have reported that a layer of ferrite separates pearlite from austenite at the growing interface (i,e., the advancing face of Fe3C is in contact with ferrite, not austenite) and the ratio of the thickness of this layer to the spacing of the pearlite has not been found to vary with alloy content or transformation temperature. This observation does not require any alteration in the above equation: however, it may require that a different significance be attached to the factor, AT. The AT term appears in Eq. [I] as a consequence of a factor for the amount of growth per complex transformed at the interface.* It is based *This is not the sole reason for expecting such a relationship. See Ref. 7._______________________________________________________ on a l/2 ?T dependence of the interlamellar spacing. Zener2 has developed a theory which leads to Soa 1/?V where So = the pearlite interlamellar spacing. This relation is as consistent with the data on interlamellar spacing as other relationships, although, as pointed out by Mehl and Hagel,1 the data are not sufficiently precise to consider the relationship as confirmed. The work of Darken, Fisher, and Galligan8 requires that the above be written where f = the thickness of ferrite layer between austenite and pearlite. The present paper is concerned with iron-carbon-manganese eutectoid alloys and reports experimental measurements of the following kinds: 1) the effect of

Jan 1, 1961

By Irving Cadoff, Kurt Komarek, Edward Miller

Barium can be purified by equilibration with titanium. The melting point of barium was found to be 726.2° i 0.5 °C. The room-temperature lattice parameters of BaTe and Bask are 7.004 * 0.002A and 6.600 * 0.002A. Melting points for BaTe and Base were found to be 1510° * 30°C and 1830° ± 50°C, respectively. HIGH-purity barium and its compounds are difficult to prepare because of the reactivity of barium with the atmosphere and the large heats of formation of the compounds. Purification of barium by vacuum distillation,' and the preparation and properties of barium oxide2 and barium sulfide3 have been reported. However, little has been done on the homologous compounds barium selenide and telluride. PURIFICATION OF BARIUM Distilled barium obtained from King Laboratories was used as the starting material. The analysis supplied with the metal showed the presence of: 0.4 wt pct Sr, 0.001 pct Mg, 0.02 pct F, 0.003 pct Cu, 0.005 pct Na and less than 5 x 10-3 wt pct of any other metallic impurity. Analyses for oxygen and nitrogen were not available. Since there is evidence4 that any barium nitride present in the starting material may decompose on distillation producing nitrogen which can contaminate the distillate, further purification was performed. At elevated temperatures, any nitrogen and oxygen present in barium should be removed by reaction with titanium. Assuming that the solubility of oxygen in liquid barium is negligible near the melting point of barium, any oxygen present will be in the form of BaO. Removal of oxygen from molten barium is expressed by the equation: BaO(S)+ TixOy(S) = Ba(l)+ TixO(y+1)(s) where Ti,Oy and TixO(y+1) are solid solutions of oxygen in titanium. At 1000°C, the change in free energy for this reaction is negative for (y+1)/x +y+1) x (100) 17.5 at. pct O.5 Since reaction with commercially pure titanium (containing 0.07 wt pct oxygen) results in a free energy change for the reaction of -19 kcal per g-atom, slight solubility of oxygen in barium would not hinder the oxygen removal. Since comparable thermodynamic data are not available to permit calculation of the partition of nitrogen between liquid barium and titanium, a similar quantitative relationship cannot be obtained. However, on the basis of work by Kubaschewski and Dench,5 complete removal of nitrogen from liquid barium can be expected. Since the melting point of barium is depressed markedly by small additions of nitrogen,' the change in melting point during reaction of barium with titanium was used to follow the purification reaction. MELTING POINT OF BARIUM A 50-g sample of barium was sealed by arc welding under argon into an all titanium crucible provided with a thermocouple well. The melting point of the sample was determined by thermal analysis, using a Pt/Pt-10 pct Rh thermocouple which was calibrated according to National Bureau of Standards specification6. The crucible was then heated for 48 hr at 950°C in vacuum and the melting point redetermined. This procedure was repeated until three successive thermal analyses agreed within ±0.5oC, the limits of error of the analysis. The melting point increased from an initial value of 720.0°C to a final value of 726.2°C. Analysis on samples quenched from 950°C gave a solubility value of 0.004 wt. pct Ti. Assuming that the titanium-barium phase diagram is similar to those of titanium-magnesium7 and titanium-calcium,8 the solubility of titanium in liquid barium decreases with decreasing temperature. Therefore, the solubility of titanium in liquid barium should be less than 0.004 wt. pctat the melting point (726oC), and the effect of dissolved titanium on the melting point would be negligible. Addition of up to 3 wt pct Sr does not significantly change the melting point of barium,7 so that the effect of the 0.4 wt pct Sr in the starting material will also be negligible. The value of 726.2" ± 0.5C obtained for the melting point of barium can be compared .with a determination carried out by Keller and coworkers in low-carbon steel crucibles,' who obtained a value of 725± 1C, using barium purified by fractional distillation. The higher value obtained in the present investigation is indicative of the effectiveness of titanium in removing traces of nitrogen. PREPARATION OF BaTe AND Base The compounds were prepared by direct reaction

Jan 1, 1961

By J. S. Kirkaldy, W. J. Slater, G. R. Mason

Activation analysis for copper and manganese in 10 mg samples machined from aluminum-base diffusion couples has been used in the study of ternary diffusion in the Al-Cu-Mn system. The accuracy, economy and convenience of the method are considered to be very competitive with other methods for analysis of such small samples. Neutron-activated aluminum-zinc alloys in the range from 10 per cent to 100 per cent zinc have been equilibrated with a known volume and the vapors subsequently condensed and collected. The vapor pressures relative to pure zinc (the thermodynamic activities) have been determined at 550°C' by recording the relative count of /'-rays from zn65 in the collected samples. The method is a sensitive one, although subject to serious obstacles to experimental accuracy.

Jan 1, 1961

By D W. MacFadgen

THE operating advantages of the longwall retreating system of mining, as opposed to longwall advancing, have been well known for many years. Although retreat walls have been successfully worked in the past in some of the collieries of the Dominion Steel and Coal Corporation in the Sydney, Picton and Cumberland areas, the adoption of longwaIl retreating as the general sys-tem of mining has not been attractive in the Sydney field, due to the time and expenditures required to drive roadways to the boundary. In addition. difficulties in maintaining effective ventilation on the faces, especially under conditions of low barometer, have made the operators in the field hesitant to adopt this as the general system of mining. The development of a machine capable of rapid entry drivage has made possible the driving of the necessary road ways to the boundary at a moderate cost, completely changing the picture and making a retreat system more attractive.

Jan 1, 1961

By R. M. Becker

THIS IS a progress report on a study of the application of statistical practice to mine and mineral sampling undertaken by the Federal Bureau of Mines to provide sound theoretical and practical data for use of the mineral engineering professions. One of the problems in mine sampling is determining the reliability of the results. For certain types of sampling data (random data) the reliability of the results can be stated mathematically. For other types of data there is no known means of quantitatively expressing reliability. In applying measures of reliability to a set of random data, some knowledge or assumption of the characteristics of the particular data is necessary. In many fields these characteristics are well known. However, in mining, where the data are in assays, these characteristics are not well known.

Jan 1, 1961

By T. A. Klinefelter

KYANITE and the related minerals sillimanite, andalusite, dumortierite, and topaz are natural aluminum silicates that can be used in production of mullite (3A1203?2Si02), a superduty refractory material. Kyanite and sillimanite are the dominant ones because of their greater abundance. The only other kyanite-group mineral presently used for mullite production is andalusite. Mullite is not an absolutely essential material as other refractories can be used in the same heat range. However, it is a highly satisfactory one. It is the only compound of silica and alumina that is stable at high temperatures. It has low thermal expansion and is very resistant to corrosive melts and to thermal shock. For these reasons mullite refractories have a much longer life than standard fire-clay refractories under the same conditions. Worldwide, excluding U.S.S.R., the refractories industry consumes about 150-000 short tons annually of kyanite group minerals and synthetic mullite.

Jan 1, 1961

By C. F. Austin

lined-cavity shaped charge is an explosive mass with a cavity at one end and the detonator at the opposite end. The cavity is lined with a dense material, such as metal, glass, or a ceramic. Such an explosive device can be a useful drilling tool. Lined-cavity shaped charges are constructed to do just one thing — create a high-velocity jet of liner particles that can bore a hole in a target. The liner must be smooth and symmetrical, and high-velocity, high-brisance explosive must be used rather than ordinary commercial dynamites. Since the depth of target penetration is related to jet length, a standoff distance should be provided to hold the charge a short distance from the target. This allows the jet to lengthen prior to target impact. The best results are obtained when the jet is perpendicular to the target surface. Charges can be positioned for firing by taping them to a stake or simply laying them on a flat surface at the proper distance and angle with respect to the target. EFFECTS ON THE TARGET When a jet strikes a rock target, the jet and target particles at the jet-target interface move outward radially with respect to the jet axis to create a tremendous compressive load on the target material immediately below the target surface. The result is a shallow crater formed by spalling at the point of target entry. After cratering ceases, the steadily deepening hole continues to form as a result of the jet's imparting radial motion to the target particles at the jet-target interface. These radially moving particles collide with one another and with the hole walls, the net motion of the spent jet particles and target particles being back out of the hole in an axial direction (the backblast). The backblast keeps the penetration free of debris during penetration, al- though loose surface rock often falls into the hole after penetration ceases. The sudden stress relief after penetration ceases sometimes results in minor hole-wall rebound, which can partially fill the hole. The hole bottom will taper as the jet velocity decreases, until the jet velocity is too low to maintain the backblast. At this point the remaining jet and the following slug material are simply packed into the hole bottom. HAZARDS INHERENT IN USE Air blast from the highly brisant explosives used in shaped charges can damage nearby structures

Jan 1, 1961

By Phillip M. Busch

VANADIUM was discovered in 1801 by Andres Manuel del Rio, a professor of mineralogy at the School of Mines in Mexico City. Del Rio called the new element erythronium. Later, he decided that this was not a new element but a form of chromium. Not until 1831 did Nils G. Sefström, a Swedish chemist, positively determine that the metal was a new element which he named vanadium, after Vanadis, the Scandanavian goddess of beauty. From 1831 to the 1860's and 1870's, the comparative scarcity and high cost of vanadium restricted its commercial use. Shortly after 1870, more information was obtained on the physical and chemical properties of the element. Thereafter, research on vanadium has dealt largely with its use as an alloying element in steel and nonferrous metals. Vanadium is a widely distributed element in the earth's crust. It is more abundant than many metals but is disseminated in small quantities throughout a great variety of rocks. Relatively few deposits have contained ore of a grade suitable to warrant development for vanadium alone. It is produced in conjunction with uranium, iron, titanium, aluminum, and lead-zinc ores.

Jan 1, 1961

By John C. Machisak

INJURY data and experience, with related employment statistics, at coal mines in the United States for 1958 are presented in this publication under the following general sections: (A) General injury experience: Tables 1 to 6: Current data on number of injuries, degree of injury, frequency and severity rates, average severity, and related employment and production data by work location. (B) Selected injury experience: Tables 7 to 38: Current data on principal agencies of injuries, distribution of injuries by part of body injured, days of disability, extent of hand and machine loading, experience by employment and production groups, and various other operating factors and physical conditions related to the occurrence of injuries at the Nation's coal mines. (C) Injury experience by States: Tables 39 to 43: Current data on injuries, their causes, rates of occurrence and severity; employment and production data by work location, by States. (D) Major disasters: Table 44: Data on major disasters from 1949 to 1958. A major disaster is defined as a single accident in which, or as a result of which, five persons or more are killed instantly or die at any time thereafter. (E) Historical coal-mine-fatality experience: Tables 45 to 49: Data on fatalities and their causes at bituminous coal, lignite, and anthracite mines from the earliest records through 1958.

Jan 1, 1961

By Lloyd L. Farnham

This paper is one of a series describing the mineral resources of the Nation, including location, history, production, ownership, geologic setting, and mining methods for most of the known manganese deposits in New Mexico, With few exceptions these descriptions are the result of field examinations by the writer. Of 32 counties, the 12 which occupy the western half of the State contain all the known manganese deposits, These counties are Catron, Dona Ana, Grant, Hidalgo, Luna, Rio Arriba, Sandoval, Santa Fe, Sierra, Socorro, San Juan, and Taos. (See fig. 1.)

Jan 1, 1961

By Bolmerm R. L.

This information circular describes the application of modern methods and equipment in sinking a vertical, 916-foot-deep by 16-foot-diameter, concrete-lined access shaft for the Harold D. Roberts tunnel near Montezuma, Colo. Although the access shaft was merely a construction facility to expedite completion of a 23-mile water-diversion tunnel, a circular shaft of this type would be equally adaptable for mine development or other purposes. Comparable sinking methods and equipment could be used for similar shafts within a wide range of sizes and depths. SUMMARY In addition to excavation and support of the basic access shaft to tunnel grade, the overall sinking project involved initial preparation of a surface site and plant, subsequent construction of a tunnel-level station and loading pocket at the shaft bottom, and ultimate conversion of the basic shaft into a hoisting and access facility for tunnel operations. The sinking period for the basic shaft approximated one year. Actual tunnel operations through the converted shaft began on April 3, 1958, approximately 19 months after the project was initiated.

Jan 1, 1961

By Roy F. Abernethy

This publication by the Federal Bureau of Mines contains the free-swelling indexes for 2,812 samples of coal received from January 26, 1948, to July 15, 1957, and the Hardgrove grindability indexes for 2,339 samples of coal received from June 23, 1947, to July 15, 1957. These data consist of the origin of the sample, bed name, and analysis. Summaries show the high, low, and average values for each bed. The free-swelling index is a relative measure of the caking properties of a coal when burned on a grate, and the Hardgrove grindability index is used to estimate the relative ease of grinding coals. Development and standardization of these test procedures were at a much later date than those for the more familiar tests, such as proximate and ultimate analyses and calorific value determinations. The newer tests are valuable in supplying supplemental information, particularly for assisting in evaluating coals used in certain types of combustion processes.4/ Data from such tests are important to designers of burning equipment, operating engineers, and purchasing agents. Data from these newer tests are scarce and scattered through a number of publications. Special projects were started to determine the free-swelling and grindability indexes, in addition to the regular analyses on the tipple samples collected by Bureau engineers. Since the start of these special projects, all the coal fields of commercial importance have been adequately covered in the sampling program.

Jan 1, 1961

By J. R. Stone, G. D. Storm

Design features and operating methods of ASARCO's new unit for the continuous melting and casting of tough pitch copper at Perth Amboy are described. Preliminary studies made for determinitzg economic feasibility and design details are also presented. Cathodes are preheated with oil and melted in an electric furnace at 40-ton per hr rate, from whence the molten copper is fed simultaneously to the wire bar casting wheel and the continuous casting of cakes and billets. In the early 1950's the copper refining furnaces at ASARCO's Perth Amboy plant reached a point where major expenditures would have to be made for replacement of the furnaces and the auxiliary equipment. The problem was whether to continue with the standard reverberatory furnaces or to consider continuous melting in electric furnaces. Reverberatory furnaces have been used for years in the copper industry and for some 50 years at the Perth Amboy plant. Reverberatory melting practice, however, imposes certain restrictions. First, labor costs are high, usually requiring three shifts per day. Secondly, they operate on a batch basis, the conventional operation being a 24-hr cycle including

Jan 1, 1961

By I. C. Edwards, W. J. Dengler, D. F. Lillie

By 1959 the milling plant of Bicr oft Uranium Mines Ltd. had been in operation for two years. During this time many changes, both physical and chemical, had been made in an effort to improve plant efficiency. Investigational work had been facilitated greatly by the fact that the Bicroft mill consists to a large extent of two parallel circuits; however, the number of interrelated variables is great, and it is difficult to measure, let alone control, all of them simultaneously. The Bicroft mill staff feels that much progress has been made, but there is much yet to be done and many questions still to be answered. The Addendum outlines noteworthy developments to the fall of 1960. The property of Bicroft Uranium Mines is located in Cardiff Township in the Highlands of Haliburton County, southeastern Ontario, some 140 miles northeast of the city of Toronto and 70 miles due north of Lake Ontario. The ore-bearing dikes are classed as mafic. nonsegregated, syenite and granite pegmatite. Microline, microline perthite, quartz, sodic plagioclase, and pyroxene or biotite are the main constituents with uraninite, uranothorite, fluorite, titanite, cryolite, calcite, pyrite, pyrrhotite, and molybdenite as important accessories. The mill was designed for a nominal capacity of 1000 tpd with all sections except grinding, leaching, and clarification capable of 1500 tons. Feed was entered to one mill unit on Oct. 29, 1956, and to the second parallel line on November 13. The total feed rate was at 1000 tpd on November 15. The first precipitate was produced on November 12. By August 1958 the average entry had been increased to 1410 tpd without the installation of any additional major equipment. The overall recovery has been greater than 93.0 pct on several months, but a recovery of 92.5 pct appears to be most economical. Operating time for eight months in 1958 has been 97.1 pct although all grinding mills were relined during this period. A generalized flowsheet is shown in Fig. 1 and the plant layout in Fig. 2. The ore is crushed in three stages to -3/8-in. and then is wet ground in two stages to approximately 50 pct -200 mesh. The pulp is thickened to 68 pct solids and agitated by air for 33 hr at a pH of 1.9 as maintained by the addition of 93 pct sulfuric acid. Separation of the uranium-bearing pregnant liquor from the residue is accomplished by five-stage countercurrent decantation (CCD), following which the liquor is clarified for removal of fine slimes. The clarified liquor is treated in the ion exchange equipment in which both the upgrading of the solution and the removal of some impurities is accomplished. Addition of magnesia to the high-grade eluate from ion exchange precipitates the uranium as sodium-magnesium diuranate. After being filtered and dried this precipitate is packaged in drums for shipment. The residue from CCD and the barren solution from ion exchange are neutralized by the addition of lime and pumped to the tailings disposal area. CRUSHING All crushing is done on surface in a conventional three-stage unit consisting of a 36x48-in. Birdsboro Buchanan jaw crusher, a 4 1/4-ft standard Symons cone crusher, and a 4-ft Symons shorthead. Both the two primary and the two secondary screens are 4X 8-ft Symons rod decks. Moisture content of the feed ranges from 3 to 4 pct. In the initial installation the primary screens were equipped with 5/16-in. rods at 1 1/4-in. spacing and the secondaries with 5/16-in. rods at 1/2-in. spacing. With this arrangement the average crushing rate was better than 180 tph. To reduce the work required of the grinding circuit as tonnage was increased, several changes were made in the crushing circuit. By September 1957, both secondary screens had been reduced to 3/8-in. spacing, and rod size had been reduced to 3/16-in. to provide increased open area. Average throughput was reduced to 160 tph. Later, 50 pct of one primary screen was decreased to 1/2-in. opening to better balance the load between the two cone crushers. In May 1958,

Jan 1, 1961

By Lotfollah Nahai

India is one of the fortunate countries that has substantial reserves of many raw materials for producing steel; however, for historical and economic reasons these resources have not been aggressively developed until recently. Production of steel on an industrial scale started in 1912, but annual output was below 1 million tons3 in 1938 and about 1,500,000 tons in 1951. Expansion of the steel industry became one of the objectives of India's economic planning after independence in 1947. The Planning Com¬mission's 1951 report emphasized the importance of steel production, but financial obstacles, emphasis on river valley projects, and priority for increased food production in the First 5-Year Plan (1951-56) did not permit large-scale expansion of steel capacity. The 1956 planned target for crude steel, only 1,650,000 tons, was actually exceeded by production of 1,737,900 tons. In the Second 5-Year Plan (1956-61), however, construction of steel plants received a high priority. Currently three new steel plants, each with an annual capacity of approximately 1 million tons of steel, are nearing completion in Rourkela, Dorgapur, and Bhilai. Simultaneously, the two privately owned steel plants and the small State-owned plant in Mysore have been enlarged and modernized. As a result, India's crude steel capacity will be about 6 million tons in 1961, almost evenly divided between Government-owned and privately owned plants. Actual production in 1960 totaled 3,254,000 tons compared with 2,434,000 tons in 1959. Assuming that the population will be 430 million and that capacity production will be achieved, the 1961 per capita steel output will be only 31 pounds compared with 200 pounds for Japan in 1959. Some Indian experts feel that development of steel-consuming industries merits greater emphasis in the Third 5-Year Plan (1961-66) than expanding existing steel-producing facilities. Nonetheless, the Third 5-Year Plan calls for a 10 million ton annual capacity by 1966 to be achieved by expanding the Government-owned plants in Bhilai, Durgapur, and Mysore, and by establishing a new integrated Government plant in Bokaro, in the State of Bihar.

Jan 1, 1961

By Alvin Kaufman

Following the gold discoveries of the late 1800's much was written regarding the vast mineral resources of Alaska, and much interest was aroused, and yet over the years little has been done to exploit those resources. Between 1906 and 1935 the Kennecott copper mines of the Copper River region were the keystone of the mineral economy (fig, 1). After the depletion of the Kennecott ore reserves and aided by the statutory rise in the price of gold, the yellow metal became the major mineral commodity. In recent years it has shared the honors with coal and sand and gravel. Although this trend should indicate the development of a stable, diversified mineral industry, yet the trend of production is downward. This report results from a study by the Federal Bureau of Mines to delineate the problems and to forecast the growth of the mineral industries in Alaska. The need for such a blueprint for the future was enhanced by the population explosion in the United States and the consequent eventual need for additional living space, as well as the current interest in new areas, and the necessity to develop reserves of strategic minerals.

Jan 1, 1961

The Emmaville district is located in the northern part of the New England region of New South Wales some 400 miles north of Sypney. Better known for its substantial production of tin and tungsten ores, the Emmaville district has also been the centre of a small and somewhat sporadic production of copper, arsenic, silver, lead and zinc. This production has come principally from three mines-Webb's Silver Mine, Webb's Consols and, to a lesser extent, Pye's Creek Silver Mine.The ores from these mixed sulphide deposits are texturally and mineralogically complex and the problems of beneficiation are rather formidable.GENERAL GEOLOGYThe main rock type of the Emmanville district is a dark, brown to blackish massive siltstone of Permian age which overlies acid volcanics possibly also of Permian age. In places the siltstone gives way to conglomerates and standstones with discontinuous lenses of calcareous sediments (David, 1887).The whole series is folded but because of the massive character of the more expansive siltstone, detailed structure is difficult to delineate.Toward the close of Permian time the area was intruded by a series of intermediate to acid plutonic and hypabyssal rocks. Mineralization within the region is ascribed to the more acid of these intrusives-granodiorite, granite, and quartz-felspar porphyry (Lawrence, 1957, 1960).No intrusive rocks outcrop in the immediate neighbourhood of Webb's Silver Mine though granite occurs two miles...

Jan 1, 1961

By S. C. Lindsay, D. J. I. Evans

With each year that passes hydrometallurgical processes are being more widely used to recover base metals from ores and concentrates. Generally these processes involve liquid-solid separation of metal-bearing liquors from barren residues. This may be done by countercurrent decantation, thickening and filtration, or filtration or centrifuging only. Choice of method is governed by chemical and physical characteristics of the pulps; cost of labor, power, materials, and waste disposal; and even by the amount of capital available when building plans are under way. Since all these controlling factors change with time and location and with technological advances, the separation method must be re-evaluated when new plants or expansion programs are contemplated. At the Sherritt Gordon refinery in Fort Saskatchewan, Alberta, an ammonia pressure leaching process recovers nickel, copper, and cobalt from the nickel concentrate produced at Lynn Lake, Manitoba.' Pregnant liquors are separated from the barren leach residues in a thickener-filter liquid-solids circuit. The decision to use this type of separation was based on a summation of economic factors during the pilot plant and design stages. Although countercurrent decantation was not incorporated in the refinery flowsheet, the problems solved in thickening the leach residues and the experience gained in operating the commercial plant are pertinent to countercurrent decantation, particularly where it is applied to slow-settling leach slurries or slurries having a relatively high vapor pressure of a valuable, obnoxious, or poisonous vapor. The nickel sulfide concentrate is leached under pressure with air and ammonia to dissolve the nickel, copper, and cobalt and most of the sulfur, leaving a residue of iron and siliceous material to be discarded.2,3 During a subsequent distillation stage to recirculate part of the ammonia used for leaching, copper is recovered from the solution as a high-grade sulfide. After further purification to remove traces of copper and undesirable sulfur compounds, the nickel is precipitated by hydrogen under pressure at an elevated temperature. The cobalt, together with unreduced nickel, is precipitated as a mixed metal sulfide by hydrogen sulfide for eventual cobalt recovery. The end solution is then evaporated to produce a crystalline ammonium sulfate byproduct. As shown in Fig. 1, the nickel concentrate is leached in a continuous, two-stage, countercurrent circuit. In the first stage, or adjustment leach, the fresh concentrate is partially leached (60 pct) in the liquor produced in the second stage, or final leach. This system is necessary to produce a solution suitable for treatment in the copper sulfide precipitation step. Both stages are carried out at 170" to 180°F under a total pressure, with air, of 100 to 120 psig in large, horizontal, water-cooled autoclaves. The autoclaves are divided by overflow weirs into four compartments to give a certain amount of staging during leaching. In each compartment a relatively slow-speed (100 to 150 rpm) agitator maintains the nickel concentrate in suspension. Adjustment leaching, carried out in two autoclaves in parallel, is followed by a thickener-filter liquid-solids separation. Final leaching is done in two parallel trains of three autoclaves using aqueous ammonia to extract metal values from the adjustment leach residues. This too is followed by a thickener-filter circuit. Here the residues are repulped once with aqueous ammonia and twice with process water and refiltered after each repulping stage, to insure a maximum recovery of soluble metal, before being discarded to the tailings pond. The adjustment leach is followed by a seven-disk Dorr-Oliver American filter, with an agitator in the filter boot to keep the residues in suspension. Agitators have also been installed beneath the four ten-disk filters following the final leach thickener. Slurry Characteristics: A typical analysis of the Lynn Lake nickel concentrate, produced by differential flotation,' is given in Table I. The concentrate comprises 35 to 49 pct pentlandite, (Fe, Ni)S; 3 to 6 pct chalcopyrite, CuFeS,; 1 to 4 pct pyrite, FeS2; and 24 to 42 pct pyrrhotite, Fe1-2S (where x varies between 0 and 0.2) together with various siliceous minerals such as actinolite, hornblende, and talc. During leaching, the sulfur in the pentlandite, chalcopyrite, and part of the pyrrhotite is successively oxidized through the thiosulfate and poly-thionate radicals to sulfate and sulfamate." Simultaneously the nickel, copper, and cobalt are converted to soluble ammine complexes and the iron in the pentlandite, chalcopyrite, and part of the pyrrhotite is oxidized to a spongy hydrated iron oxide. The hydrated iron oxide is readily broken down by the agitation needed to sustain the concentrate in suspension and to maintain a satisfactory mass transfer of the oxygen from the air to the slurry. Table I shows typical chemical, screen, and

Jan 1, 1961

By C. W. DeLong, L. L. Farnham, L. A. Stewart

This paper is one of a series covering mineral resources of the Nation . It describes briefly most of the known manganese deposits of eastern Arizona , which for the purpose of this report includes the counties of Navajo , Apache , Gila, Graham, Greenlee , Pinal , Pima , Santa Cruz , and Cochise . With few exceptions , these descriptions are the result of field examinations by the writers . Work on this paper was begun in May 1956 , and in the following 18 months 151 deposits were examined . The distribution of the deposits by counties is as follows : 1 in Navajo , and 1 in Apache , 37 in Gila , 4 in Graham, 13 in Greenlee , 25 in Pinal , 13 in Pima , 12 in Santa Cruz , and 45 in Cochise . In surveyed areas the locations of the deposits are described by section , township , and range and in unsurveyed areas , by approximate projections of these subdivisions . The township and range numbers refer to the Gila and Salt River base and meridian . The history, production, ownership , geologic setting , and mining methods used are discussed for most of the deposits that were visited . Probably , some occurrences have been overlooked , but all of the principal ones have been examined . Ore dressing studies were made of typical ores from several of the larger districts . The work was performed in the laboratory of the Bureau of Mines Tucson Metallurgy Research Laboratory . The results of these studies are given with the descriptions of the deposits from which the samples were taken .

Jan 1, 1961