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US 4,183,794 - Electrowinning zinc from a zinc sulfate solution produced by leaching zinc ore with sulfuric acid. The leach liquor is subjected to electrolysis in a first cell having an insoluble anode to deposit primary zinc on the common electrode. The coated electrode is transferred to a second cell having an alkaline electrolyte, and electrolysis is effected in the second cell to transfer zinc from the common electrode to the second cell cathode in a spongelike form which is readily stripped. R.H.C.L. Ross. Jan 15, 1980. 10 pp. (204-119). Filed: 12-22-77 (863,190). Same: Canadian 1,051,820, dated Apr. 3, 1979.

Jan 1, 1982

By O. J. C. Runnals

The intermetallic compounds NpAl2, NpAl3, and NpAl have been prepared, and examined by X-ray diffraction methods. The compounds are isostructural with the corresponding U-Al compounds. NpAl3 is face-centered cubic with a = 7.785A and has the MgCu2 structure. NpAl is simple cubic with a = 4.2624 and has the AuCU3 structure. NpAl4 is body-centered orthorhombic with a = 4.428, b = 6.268, and c = 13.71A. The published description for the atomic positions in UAl, applies also to NpAl4 THE U-Al system has been examined by Gordon and Kaufmann.1 They identified the following intermetallic compounds: UAI2, UAl3, and a third, tentatively identified as UAl5. A structure analysis of the latter compound by Borie2 proved that the formula UAl4 was more consistent with the X-ray data. The crystal structures of UAl2 and UA have been reported by Rundle and Wilson.3 UAl2 was found to be isostructural with MgCu. The atomic positions in the simple cubic UA1, cell were found similar to those in the AuCu3 structure. Three analogous compounds exist in the binary system Np-Al. X-ray data for the compounds NpAl2, NpAl3, and NpAl4 are reported below. The neptunium used in the investigation was supplied through the courtesy of the Argonne National Laboratory as neptunium (V) nitrate. It was stated to have a radioactive content of 99.25 pct Np237. The impurity limits had been established by a spectrographic analysis given in Table I. Any titanium, boron, or bismuth impurity should be effectively removed during the fluorination of NpO2 to NpF4, and sodium, lithium, magnesium or strontium during the alloying procedure. The oxalate ignition technique, described by Westrum for the preparation of finely-divided reactive PuO2, was used.The neptunium (V) nitrate solution, containing 100 mg Np, was evaporated to dryness on oxalic acid crystals. The resulting oxalate was calcined to NpO, by heating in platinum at 650°C for 2 hr. An equimolar mixture of anhydrous hydrogen fluoride and purified hydrogen was passed over the NpO2 for 2 hr at 500°C. According to Fried and Davidson" the above fluorination conditions should produce the purple-black fluoride, NpF3 However, the light-green product showed only NpF4 lines on an X-ray diffraction pattern and contained only an estimated 5 pct of finely dispersed black NpF3 particles on microscopic examination. NpAl3: An intimate mixture of 50 mg of neptunium fluoride and aluminum powder (99.99 pct pure, —28 to $60 mesh) was heated to 900°C in a BeO crucible in vacuum by a tungsten coil for 2 hr, to allow the following reaction to proceed: NPF4 + 7Al + NpAl, + 4AlF. [1] The formation and sublimation of aluminum mono-fluoride in this system was verified from an X-ray diffraction analysis of the sublimate. The diffraction pattern was similar to one obtained in the aluminum reduction of UF4 to UF3,6 showing both Al and AlF3 lines. The resulting alloy was sintered at 1100°C for 1 hr. A powdered specimen of the crystalline product was sealed in a thin-walled quartz capillary for X-ray examination. The photograph was taken on a 14.32 cm General Electric powder diffraction camera, using filtered copper radiation. The films were calibrated in terms of a NaCl standard pattern. The wavelengths used in the calculations were 1.5418A for Cu K and 1.5405A for Cu K The powder diffraction pattern showed that NpAl is simple cubic with a = 4.262 % 0.005A, and is isostructural with UA13 where a = 4.27A1 or a =

Jan 1, 1954

By John Birkinbine

(Wilkes-Barre Meeting, June, 1911). MODERN advances in practically all lines of industrial development have occurred in such rapid succession, and have been accepted so readily as accomplished facts, that a retrospect surprises us, by showing how comparatively few of the acknowledged factors of improved conditions may be considered as old. While these advances have not been confined to any country, they have been more pronounced in some than in others, and nowhere more so than in the United States, the population of which, having multiplied nearly three-fold between 1870 and 1910, demanded a proportionately greater increase in materials, supplies, and manufactured products. It therefore appears desirable to discuss mainly conditions in the United States, as concrete evidence of industrial progress throughout the world. Looking backward for but three generations, we may trace the introduction and development of canal- and steamboat-navigation; railroad-transportation; artificial illumination beyond that furnished by candles and lard-oil; quick communication by mail, and subsequently by telegraph or telephone; the manufacture of iron, beyond forms of small dimensions; the production and utilization of steel in large quantities; the economic use of mineral fuel, oil and gas, etc. The practical coincidence of the fortieth anniversary and the 100th technical meeting of the American Institute of Mining Engineers, offers temptation to recall and compare the conditions of mining and metallurgy in or about the years 1871 and 1911. A complete resume would cover phenomenal changes in mining-methods and equipment by which the output of individual exploitation has grown from scores to hundreds and even thousands of units in equal time-intervals. Extension of operations in depth and area, demanding machinery of great power

Aug 1, 1911

By C. R. Masson, M. L. Pearce

The nitrogen contents of seven specimens of iron and steel were determined by the three methods. Good agreement won generally observed between the results of the isotope -dilution and Kjeldahl determinations, tltzts establishing the reliability of the isotope -dilutzon technique. For some specimens, the vacuum-fusion values at 1650 °C were lower than the corresponding Kjeldahl values by up to 0.0015 pct absolute. These differences were eliminated for a specimen of "pure" iron by performing the vacuum-fusion analyses at 2100 ° to 2240 °C. The reason for this relatively high temperature required m the vaczium-fusion analyses is discussed. ThE determination of nitrogen in iron and steel has been the subject of numerous investigations and a concise review of the extensive literature in this field has recently appeared.' Two methods of analysis are most commonly employed, the vacuum -fusion and the chemical or Kjeldahl method. Except in special cases where the presence of certain alloying elements may present difficulties, both methods are generally regarded as applicable, although the vacuum-fusion method is more sensitive and is often preferred for samples of very low-nitrogen content. The method of isotope dilution, first employed for the determination of gases in metals by Kirshenbaum and Grosse,2 has recently been applied to the determination of nitrogen in iron and steel.3,4 The method is attractive in that it does not require the quantitative extraction of gas from the metal specimen. Staley and Svec3 modified the standard Kjeldahlpro-cedure by introducing, along with the steel sample, a measured amount of ammonium sulfate enriched in the isotope 15N. The nitrogen content was obtained by measuring the isotopic ratio 15N/14N in the resultant solution. The technique involved the distillation of nitrogen as ammonia from alkaline solution and subsequent oxidation of NH4+ in the distillate to gaseous nitrogen. The results for iron and steel samples of commercial origin were often in reasonable agreement with values quoted by the suppliers, although occasional marked discrepancies were observed, particularly with the results of vacuum-fusion analyses. Preliminary work in this laboratory4 on the determination of nitrogen by isotope-dilution in the gas phase yielded results for three samples which were significantly higher than the values obtained by vacuum-fusion. The present paper describes an extension of this work to a wider variety of samples. Analytical results by the Kjeldahl method are also reported. EXPERIMENTAL The technique employed for the isotope-dilution determinations was similar to that described elsewhere.4 Several minor modifications were made to the apparatus. These included the use of molybdenum instead of platinum-rhodium suspending wires when crucibles were degassed at temperatures above 1600°C. The circulatory system used previously for converting CO to CO2 was replaced by a single chamber, the walls of which were cooled directly in liquid nitrogen. Temperatures were measured with an optical pyrometer which had been recently calibrated and, as in the previous study, were minimum values obtained by sighting on the center bottom of the crucible. Temperatures obtained by sighting on the interior walls were always higher by 100' to 130°C with the design of crucible employed. Flat-bottomed crucibles exhibited an even larger temperature difference. To check the efficiency with which crucibles were degassed in the isotope-dilution apparatus, several experiments were done in the vacuum-fusion apparatus, suitably modified4 for isotope-dilution studies. Faster pumping speeds and higher degassing temperatures were possible, the latter due to the use of a generator with a higher power output (10 Kva at 398 kc per sec). In these experiments the combustion of CO was omitted and a correction4 was made for the contribution of CO to the ion current at m/e 28, used in determining the isotopic ratio R = l5NP4N. This correction did not seriously influence the re-

Jan 1, 1962

By Carleton C. Long, George F. Weaton

Two years ago the general features of the St. Joseph Lead Company's zine-smelting process were described.' At that time the discussion was limited to a description of the production of high-purity electrothermic zinc oxide. The present paper will describe the commercial production of electrothermic zinc metal. Since a fairly complete description of the plant of the Josephtowlz smelter was given in the earlier paper, the brief recapitulation given here will be limited to mention of additional equip-ment recently installed and a bare outline of the preparatory process. The raw material is zinc sulphide concentrate received from the company's mines in St. Lawrence County, New York. An average dry-weight analysis of this material is: Zn, 58.28 per cent; Pb, 0.56; insoluble, 1.6; SiO², 1.12; Fe, 6.4; CaO, 0.47; MgO, 0.27; S, 31.9; Cd, 0.1; Cu, 0.06; Mn, 0.24. Moisture content is usually about 3 per cent. Three 21-ft. 6-in. diameter, 12-hearth Nichols-Herreshoff furnaces roast the coacentrate to a calcine of average conlposition (first 10 months of 1938): Zn, 67.69 per cent; Pb, 0.06; insoluble, 1.75; SiO², 1.26; Fe, 7.24; CaO, 0.47; MgO, 0.36; S, 1.88; Cd, 0.1; Cu, 0.07'; Mn, 0.27. The sulphur dioxide gas, after passing through a waste-heat boiler, proceeds through a recently redesigned low-resistance gas-purification assembly to the contact sulphuric acid plant, of which the capacity is 200 tons of concentrates per day. The calcine is mixed with a portion of returned furnace residues and sintered on 42-in, by 44-ft. Dwight-Lloyd sintering machines, gas-fired. Wind-box gases pass through air coolers before filtering in a 1440-bag Dracco unit, and thence to stack. The Dracco fume is mixed with Cottrell dust, from roaster gas cleaning, and sent to the leach plant, where it is roasted (SO² to acid plant) and leached with sulphuric acid to recover, as end products, lead sulphate, zinc sulphate and cadmium.

Jan 1, 1943

By George F. Weaton, Carleton C. Long

Two years ago the general features of the St. Joseph Lead Company's zine-smelting process were described.' At that time the discussion was limited to a description of the production of high-purity electrothermic zinc oxide. The present paper will describe the commercial production of electrothermic zinc metal. Since a fairly complete description of the plant of the Josephtowlz smelter was given in the earlier paper, the brief recapitulation given here will be limited to mention of additional equip-ment recently installed and a bare outline of the preparatory process. The raw material is zinc sulphide concentrate received from the company's mines in St. Lawrence County, New York. An average dry-weight analysis of this material is: Zn, 58.28 per cent; Pb, 0.56; insoluble, 1.6; SiO², 1.12; Fe, 6.4; CaO, 0.47; MgO, 0.27; S, 31.9; Cd, 0.1; Cu, 0.06; Mn, 0.24. Moisture content is usually about 3 per cent. Three 21-ft. 6-in. diameter, 12-hearth Nichols-Herreshoff furnaces roast the coacentrate to a calcine of average conlposition (first 10 months of 1938): Zn, 67.69 per cent; Pb, 0.06; insoluble, 1.75; SiO², 1.26; Fe, 7.24; CaO, 0.47; MgO, 0.36; S, 1.88; Cd, 0.1; Cu, 0.07'; Mn, 0.27. The sulphur dioxide gas, after passing through a waste-heat boiler, proceeds through a recently redesigned low-resistance gas-purification assembly to the contact sulphuric acid plant, of which the capacity is 200 tons of concentrates per day. The calcine is mixed with a portion of returned furnace residues and sintered on 42-in, by 44-ft. Dwight-Lloyd sintering machines, gas-fired. Wind-box gases pass through air coolers before filtering in a 1440-bag Dracco unit, and thence to stack. The Dracco fume is mixed with Cottrell dust, from roaster gas cleaning, and sent to the leach plant, where it is roasted (SO² to acid plant) and leached with sulphuric acid to recover, as end products, lead sulphate, zinc sulphate and cadmium.

Jan 1, 1943

Adams, Benjamin C., Jr., Student, Univ. of Oklahoma Norman, Okla. '36 Adams, Ernest C., Student, Univ. of Illinois 908 W. Green St., Urbana, Ill. '35 Adams, George H., Student, Colorado School of Mines Golden, Colo. '36 Adams, Robert Keller Mitchell Hotel, No. 18, Golden, Colo. '35 Afshar, Hadji Khan A., Student Box 63, Univ. Sta., Moscow, Idaho. '36 Ahrenholz, Herman William, Jr., Student, Lehigh Univ Bethlehem, Pa. '35 Alger, Robert P., Student, Missouri School of Mines & Met.Rolla, Mo. '36 Allen, Paul W., Student, Mass. Inst. of Tech Cambridge, Mass. '35 Allen, R. M., Jr., Student, Virginia Polytechnic Inst., "G" Co.Blacksburg, Va. '35 Alley, John Arnold, Student, Montana School of Mines Butte, Mont. '36 Amberger, R. H., Student, Michigan College of Min. & Tech Houghton, Mich. '35 Ambrosiani, F. Peter, Student, Michigan College of Min. & Tech Houghton, Mich. '36 Andersen, Burton Lee, Student, School of Mines, Univ. of Washington Seattle, Wash. '36 Anderson, Carlos, Student, Univ. of Oklahoma Norman, Okla. '36 Anderson, J. Wendell, Student, Pennsylvania State College State College, Pa. '36 Anderson, Joseph S. Cornelia Hotel, Ajo, Ariz. '35 Anderson, R. C., Student, Univ. of Washington Seattle, Wash. '35 Anderson, Roy A., Student, Washington State College Pullman, Wash. '35 Anderson, Willard. Student. Univ. of Oklahoma Norman, Okla. '36

Jan 1, 1936

By W. W. Coblentz

THE constants in question pertain to the total radiation and the spectral radiation of a uniformly heated enclosure, or so-called black body. These constants have been determined for the range within which temperatures can be measured with thermocouples. Conversely, using these constants and suitable instruments, such as an optical pyrometer, for example, it is possible to determine temperatures far above the range attainable with the most refractory thermocouples. The formula of Stefan Boltzmann for expressing the total radiation of a black body is R = kT4 in which k is the coefficient or "constant" under discussion. In view of the fact that total radiation pyrometers are usually calibrated on an arbitrary scale, there is no great demand for an exact value of the coefficient of total radiation in absolute value. However, it is intimately connected with the constant of spectral radiation; hence, an accurate determination of the constant k gives a check on the constant of spectral radiation. The distribution of radiation in the spectrum of a black body is represented by Planck's equation, E? = C1?-5 (e c/?T - 1)-'. The spectral radiation constant c, in this formula, is useful in optical pyrometry and in establishing a high temperature scale. The numerical value of the constant c has been determined in the range of temperatures measurable with thermocouples and, also, by extending the temperature scale to higher temperatures by means of total radiation measurements.

Jan 8, 1919

By Peter Allen, John R. Stuermer

COUNTRY RISK ANALYSIS A company operating in a foreign country assumes all the risks that it would at home. However, beyond these, it assumes risks that arise from the unique political economic, financial conditions of the foreign country which it operates. These risks are known as "country risks". The importance of these risks has grown appreciably in the recent past (particularly in the aftermath of Iran's revolution in the late 1970s and Mexico's debt rescheduling of 1982-83), and as a result, country risk analysis has become an increasingly useful management tool. Country risk analysis seeks to assess the probability that countries will service their aggregate foreign liabilities on time, as well as the probability that country conditions could inhibit the debt service capability of private sector debtors. This involves assessment of both the ability and the willingness of countries to meet their external liabilities in a timely fashion. "Ability" is linked economic and financial variables: principally, a country's economic structure, foreign exchange generating capacity and debt service burden. On the other hand, "willingness" is tied to political factors: mainly the economic/financial orientation of government policy and the stability of economic decision making. Ability: Economic, Financial Conditions and Prospects Judgement of a country's ability to meet foreign financial obligations begins with a detailed study of a country's external balance sheet which compares the country's foreign liabilities and assets. Table 1 below presents country balance sheets for Mexico, Korea and Indonesia. [TABLE 1 The External Balance 3-,-et 1983 (billions of dollars; Indonesia Mexico Korea Foreign Liabilities 31.1 67.6 40.5 By Source: Bank 13.1 75.5 26.3 Non-bank 18.0 12.114.2 Foreign Reserves 9.3 4.7 7.0] The large magnitude of Mexico's external debt (nearly $90 billion), compared to foreign assets of approximately $5 billion, illustrates the weakness of this country's external payments position, and the main reason that it has been unable to meet interest and principal payments and been forced to reschedule its foreign debt. On the other hand, the relationship between Korea's external liabilities and assets ($50.5 billion and $7 billion, respectively), is more favorable and helps explain why Korea has been able to meet debt service payments and maintain its access to the international capital markets. Indonesia has even a more favorable relationship between its foreign assets and liabilities, $31 billion and $9 billion, respectively, and its debt is mostly lower interest rate, non-bank debt, making for a less onerous debt service burden. After the balance sheet, the analysis turns to a comparison of the country's foreign exchange earning capacity and its prospective financing requirements, that is, the country's future earnings stream

Jan 1, 1985

By E. Orowan

LONG before the role of dislocations in the plastic deformation of crystals was recognized, the stress-strain field around dislocations received considerable attention in the theory of elasticity. It seems that Weingarten1 was the first to consider dislocations as simple and typical sources of internal stress in elastic solids, and Volterra2 and Timpe3 investigated their elastic field. Fig 3-1 shows the two basic types of dislocations considered by Volterra. The first arises by making a straight-edged .plane cut in a solid, displacing the two surfaces of the cut in their own plane relatively to one another by a constant amount in a direction perpendicular to the edge, and welding them together; this is shown in Fig 3-la, where the outer cylindrical surface represents the boundary of the solid, and the thin spiral cylinder around the edge of the cut has been removed in order to avoid infinities of stress and strain. The second type of dislocation is shown in Fig 3-lb; here the surfaces of the cut have been displaced relatively to one another by a constant amount parallel to the edge, and then welded together. After J. M. Burgers,4 these two types are called edge and screw dislocations, respectively. They represent simple limiting cases; in the general case, the surfaces of the cut are displaced by a constant vector that lies in the plane of the surface but has components both parallel and perpendicular to the edge of the cut. This vector is the Burgers or slip vector. The name “dislocation” was given by A. E. H. Love;5 it was bor-

Jan 1, 1954

By George Weaton

Two years ago the general features of the St. Joseph Lead Company's zinc-smelting process were described.1 At that time the discussion was limited to a description of the production of high-purity electrothermic zinc oxide. The present paper will describe the commercial production of electrothermic zinc metal. Since a fairly complete description of the plant of the Josephtown smelter was given in the earlier paper, the brief recapitulation given here will be limited to mention of additional equip-ment recently installed and a bare outline of the preparatory process. The raw material is zinc sulphide concentrate received from the com-pany's mines in St. Lawrence County, New York. An average dry-weight analysis of this material is: Zn, 58.28 per cent; Pb, 0.56; insoluble, 1.6; Si02, 1.12; Fe, 6.4; CaO, 0.47; MgO, 0.27; S, 31.9; Cd, 0.1; Cu, 0.06; Mn, 0.24. Moisture content is usually about 3 per cent. Three 21-ft. 6-in. diameter, 12-hearth Nichols-Herreshoff furnaces roast the concen-trate to a calcine of average composition (first 10 months of 1938) : Zn, 67.69 per cent; Pb, 0.06; insoluble, 1.75; Si02, 1.26; Fe, 7.24; CaO, 0.47; MgO, 0.36; S, 1.88; Cd, 0.1; Cu, 0.07; Mn, 0.27. The sulphur dioxide gas, after passing through a waste-heat boiler, proceeds through a recently redesigned low-resistance gas-purification assembly to the contact sul-phuric acid plant, of which the capacity is 200 tons of concentrates per day. The calcine is mixed with a portion of returned furnace residues and sintered on 42-in. by 44-ft. Dwight-Lloyd sintering machines, gas-fired. Wind-box gases pass through air coolers before filtering in a 1440-bag Dracco unit, and thence to stack. The Dracco fume is mixed with Cottrell dust, from roaster gas cleaning, and sent to the leach plant, where it is roasted (SO2 to acid plant) and leached with sulphuric acid to recover, as end products, lead sulphate, zinc sulphate and cadmium.

Jan 1, 1939

By H. B. Pulsifer

ABOUT 15 varieties, or modifications, of the best magnesium available were prepared and subjected to etching tests, then examined for microstructure. Of the 30-odd etching reagents that were tried, nearly half, mostly ammonium salts, etched the metal satisfactorily. The surfacing and etching of magnesium is shown to be a very simple and quick operation. The density and hardness of magnesium were determined. The most interesting new observations relate to the finding of the hexagonal etch figures, the crystal laminations, and the fact that the metal is plastic, cold, when restrained or quickly deformed. Under slowly applied pressure cold metal deforms slightly, then shears to fracture without plastic flow. The chief structural features of magnesium are presented in two reduced photographs and 30 photomicrographs. MATERIALS TESTED There are only two sources in the United States from which new metal can be procured: the American Magnesium Corpn. and the Dow Chemical Co. No pronounced structural or property differences in the metals from these two companies were disclosed by the work of this investigation. The following materials were obtained from the manufacturers: 1. Massive crystals of distilled metal (American Magnesium Corpn.). 2. Rods of hot-extruded metal, ½ -in. squares and 5/8-in. rounds (American Magnesium Corpn.). 3. Sheet magnesium, 0.005-in. thick (American Magnesium Corpn.). 4. Cast stick metal, 1 3/8-in. dia. (Dow Chemical Co.). 5. Magnesium-aluminum alloy, hot-rolled plate, ½ in. thick (American Magnesium Corpn.). From these materials, were prepared: 6. Sections from a furnace-cooled ingot made from distilled crystal. 7. Cold-strained pieces from (6), (4) and (2). 8. Pieces cold-squeezed to fracture from (6), (4) and (2). 9. Hammer-struck pieces from (6), (4) and (2). 10. Steam-hammer struck pieces from (6), (4) and (2). 11. Sections from furnace-cooled ingot of the magnesium-aluminum alloy.

Jan 1, 1928

By H. B. Pulsifer

.ABOut 1.5 varieties, or tnodifications, of the best rnagnesiurn available were prepared and subjected to etching tests, then examined for micro-structure. Of the 30-udd etching reagents that were tried, nearly half, mostly ammonium salts, etched the metal satisfactorily. The surfacing and etching of magnesium is shown to be a very simple and quick operation. The density and hardness of magnesium were determined. The most interesting new observations relate to the finding of the hexagonal etch figures, the crystal laminations, and the fact that the metal is plastic, cold, when restrained or quickly deformed. Under slowly applied pressule cold metal deforms slightly, then shears to fracture without plastic flow. The chief structural features of magnesium arc presented in two reduced photographs and 30 photomicrographs. Materials Tested There are only two sources in the United States from which new metal can be procured: the American Magnesium Corpn. and the Dow Chemical Co. No pronounced structural or property differences in the metals from these two companies were disclosed by the work of this investigation. The following materials were obtained from the manufacturers: 1. Massive crystals of distilled metal (American Magnesium Corpn.). 2. Rods of hot-extruded metal, ½-in. squares and 5/8-in. rounds (American Magnesium Corpn.). 3. Sheet magnesium, 0.005-in. thick (American Magnesium Corpn.). 4. Cast stick metal, 13/8-in. dia. (Dow Chemical Co.). 5. Magnesium-aluminum alloy, hot-rolled plate, ½-in. thick (American Magnesium Corpn.). From these materials, were prepared: 6. Sections from a furnace-cooled ingot made from clistilled crystal. 7. Cold-strained pieces from (6), (4) and (2). 8. Pieces cold-squeezed to fracture from (6), (4) and (2). 9. Hammer-struck pieces from (6), (4) and (2). 10. Steam-hammer struck pieces from (6), (4) and (2). 11. Sections from furnace-cooled ingot of the magnesium-aluminum alloy.

Jan 1, 1880

By R. C. Ferguson, Harlowe Hardinge

Oscar Johnson—In my opinion, the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. Comparing the entire groups of operators with those who have had the opportunity to make slow-speed mill studies, I think you will find the latter small in numbers. Most managers want the equipment worked to its maximum output. There are, however, some installations where plant and mill sizes are such that they can do the job with reduction of mill barrel speeds. The past and the present installations of the industry are laid out to get the most capacity for the least capital outlay. This is the case even with the plants of Chile Exploration, International Nickel, Morocco, and Anaconda, now under construction or being changed. The industry recognizes that most all equipment it buys today is good and can be depended upon for efficient performance. Under this scheme of things, I am doubtful that slow-speed ball mill operation will be generally applicable. With reference to the U. S. Bureau of Mines laboratory tests, I think table II could have been omitted. It is inconclusive as to maximum efficiency for the low-pulp level mill on hard ore. There should be no question about this point. However, data on mill speeds can be found to substantiate various theories as well as refute them. Gow, Guggenheim, Campbell and Coghill, in their paper on Ball Milling,' believe their 2 x 2 ft laboratory mill reflects results that can be expected from large mills. If so, then referring to their table 11, they state, "The conclusion to be drawn from this second series is that high speed, not exceeding 72 pct of the critical, favors capacity, as before, but that with proper conditions of operation high speeds may give as good efficiency values as low speeds. In this case the efficiency values are nearly constant. A horizontal curve would indicate that the amount of grinding was directly proportional to the power expended, and these tests suggest that such a coildition can be made to exist in commercial operations." Table II (From Paper by Gow et a1)2 Speed. Pot Critical 32 42 52 62 72 82 Capacity: Surface tons per hr (65- mesh) 266 42.1 54.4 65.9 74.3 74.1 Surface tons per hr (200- mesh) 56.1 87.4 112.7 137.1 154.2 153.0 Efficiency: Surface tons per net hp hr (65-mesh) 35.7 36.3 36.3 35.4 34.3 32.3 Surface tons per net hp hr (200-mesh) 75.3 75.3 75.1 73.7 71.0 66.0 Ore in mill, 1.b. 98 100 100 113 122 165 The field performance data, table 111, represents much effort in its collection and preparation. But, one must realize that there are many variables that effect the efficiency of grinding mill operation, and too much must not be assumed as to the effect of some specific change. Possibly with changes in mill speed, the results might be more consistent by also a change in ball rationing, type of ball, volume of ball charge,. p.ulp level and amount of pulp in the mill, pulp consisting, design of liner, circulating load, etc. Also, changes in ore character must be reckoned with when evaluating grinding performance. At present the Climax Molybdenum Corp. is running at much reduced capacity. Mr. James Duggan informs me that at mill speeds of 17 rpm, they save a $0.025 per ton on liners and $0.025 per ton in power, but, if the demand for molybdenum increased, he would go back to higher speed to obtain maximum tonnage, as the values from the increased tonnage would far more than offset the one half saving at the slower speed. The Jnspiration ran a six months' test between mills running 21 rpm and 23.5 rpm. The slower mills ground 10 pct less ore with a slight saving per ton, but when the reduced plant tonnage was checked back into the actual cost figures of concentration, the high-speed mills with their greater tonnage showed considerable advantage. To be convinced of possible practical results from the predictions in the conclusions, I think we would have to rely on the analysis of expert cost accountants to furnish the necessary proof figures. Hardinge and Ferguson are to be commended for the work in preparing this paper. I am convinced that our Massco engineers should go into higher speeds with our equipment. Harlowe Hardinge (authors' reply)—For one, I heartily agree with Mr. Johnson's opening statement that the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. It was on this basis and for this reason the paper was written. Mr. Johnson, on the other hand, takes the position that, on the whole, low speeds are not justified from the economic standpoint, basing his principal reason on the fact that lower mill speeds cut mill capacities and hence reduce the gross income from the product produced. There is no denying this point. It is almost axiomatic. It is for this very reason that the overall advantage of lower mill speeds has been discounted and even overlooked. It was for this reason mainly that the paper was written in the first place. It is one thing to plan an efficient operation at the outset, basing one's figures on the tonnage requirements at the time, and it is quite another to be confronted with the problem of increasing the output of an existing installation at a minimum of capital expenditure. Economic consideration of a new installation is greatly influenced by referring to an old one. Too often, the analyst assumes that if this practice is followed in the new installation, one would not go wrong. It is just here that he may be wrong. Past practice and low capital expenditure are all too frequently given priority over the engineer's analysis of operating costs. When we are able to start fresh, we should give proper weight to other economic factors which do not exist in an old installation. It is these economic factors that make it possible to spend at the outset just a little more money and get it back in a matter of months and effect big savings for years to come. F. C. Bond—This paper is of considerable importance in that it emphasizes a modern trend to operate ball mills at somewhat slower speeds than formerly. We have checked the data in the paper with that obtained

Jan 1, 1951

By Harlowe Hardinge, R. C. Ferguson

Oscar Johnson—In my opinion, the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. Comparing the entire groups of operators with those who have had the opportunity to make slow-speed mill studies, I think you will find the latter small in numbers. Most managers want the equipment worked to its maximum output. There are, however, some installations where plant and mill sizes are such that they can do the job with reduction of mill barrel speeds. The past and the present installations of the industry are laid out to get the most capacity for the least capital outlay. This is the case even with the plants of Chile Exploration, International Nickel, Morocco, and Anaconda, now under construction or being changed. The industry recognizes that most all equipment it buys today is good and can be depended upon for efficient performance. Under this scheme of things, I am doubtful that slow-speed ball mill operation will be generally applicable. With reference to the U. S. Bureau of Mines laboratory tests, I think table II could have been omitted. It is inconclusive as to maximum efficiency for the low-pulp level mill on hard ore. There should be no question about this point. However, data on mill speeds can be found to substantiate various theories as well as refute them. Gow, Guggenheim, Campbell and Coghill, in their paper on Ball Milling,' believe their 2 x 2 ft laboratory mill reflects results that can be expected from large mills. If so, then referring to their table 11, they state, "The conclusion to be drawn from this second series is that high speed, not exceeding 72 pct of the critical, favors capacity, as before, but that with proper conditions of operation high speeds may give as good efficiency values as low speeds. In this case the efficiency values are nearly constant. A horizontal curve would indicate that the amount of grinding was directly proportional to the power expended, and these tests suggest that such a coildition can be made to exist in commercial operations." Table II (From Paper by Gow et a1)2 Speed. Pot Critical 32 42 52 62 72 82 Capacity: Surface tons per hr (65- mesh) 266 42.1 54.4 65.9 74.3 74.1 Surface tons per hr (200- mesh) 56.1 87.4 112.7 137.1 154.2 153.0 Efficiency: Surface tons per net hp hr (65-mesh) 35.7 36.3 36.3 35.4 34.3 32.3 Surface tons per net hp hr (200-mesh) 75.3 75.3 75.1 73.7 71.0 66.0 Ore in mill, 1.b. 98 100 100 113 122 165 The field performance data, table 111, represents much effort in its collection and preparation. But, one must realize that there are many variables that effect the efficiency of grinding mill operation, and too much must not be assumed as to the effect of some specific change. Possibly with changes in mill speed, the results might be more consistent by also a change in ball rationing, type of ball, volume of ball charge,. p.ulp level and amount of pulp in the mill, pulp consisting, design of liner, circulating load, etc. Also, changes in ore character must be reckoned with when evaluating grinding performance. At present the Climax Molybdenum Corp. is running at much reduced capacity. Mr. James Duggan informs me that at mill speeds of 17 rpm, they save a $0.025 per ton on liners and $0.025 per ton in power, but, if the demand for molybdenum increased, he would go back to higher speed to obtain maximum tonnage, as the values from the increased tonnage would far more than offset the one half saving at the slower speed. The Jnspiration ran a six months' test between mills running 21 rpm and 23.5 rpm. The slower mills ground 10 pct less ore with a slight saving per ton, but when the reduced plant tonnage was checked back into the actual cost figures of concentration, the high-speed mills with their greater tonnage showed considerable advantage. To be convinced of possible practical results from the predictions in the conclusions, I think we would have to rely on the analysis of expert cost accountants to furnish the necessary proof figures. Hardinge and Ferguson are to be commended for the work in preparing this paper. I am convinced that our Massco engineers should go into higher speeds with our equipment. Harlowe Hardinge (authors' reply)—For one, I heartily agree with Mr. Johnson's opening statement that the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. It was on this basis and for this reason the paper was written. Mr. Johnson, on the other hand, takes the position that, on the whole, low speeds are not justified from the economic standpoint, basing his principal reason on the fact that lower mill speeds cut mill capacities and hence reduce the gross income from the product produced. There is no denying this point. It is almost axiomatic. It is for this very reason that the overall advantage of lower mill speeds has been discounted and even overlooked. It was for this reason mainly that the paper was written in the first place. It is one thing to plan an efficient operation at the outset, basing one's figures on the tonnage requirements at the time, and it is quite another to be confronted with the problem of increasing the output of an existing installation at a minimum of capital expenditure. Economic consideration of a new installation is greatly influenced by referring to an old one. Too often, the analyst assumes that if this practice is followed in the new installation, one would not go wrong. It is just here that he may be wrong. Past practice and low capital expenditure are all too frequently given priority over the engineer's analysis of operating costs. When we are able to start fresh, we should give proper weight to other economic factors which do not exist in an old installation. It is these economic factors that make it possible to spend at the outset just a little more money and get it back in a matter of months and effect big savings for years to come. F. C. Bond—This paper is of considerable importance in that it emphasizes a modern trend to operate ball mills at somewhat slower speeds than formerly. We have checked the data in the paper with that obtained

Jan 1, 1951

By J. Kittl, D. Arias

THE decomposition on cooling of the high-temperature ß bcc phases in copper- or silver-based binary systems usually takes place by a martensitic. massive, bainitic, or pearlitic reaction depending upon the particular conditions of each system. The AgAl system is one of these systems in which a massive or a martensitic transformation can occur upon cooling from the 3 phase field. Recently Hawbolt and Brown1 observed both massive and martensitic transformation products in a 22.7 at. pct A1 alloy. In the present work the massive and the martensitic transformations were investigated in a number of alloys for the composition range between 23.2 and 27.9 at. pct Al. The alloys were prepared using high-purity silver, 99.999 pct, supplied by ASARCO and aluminum, 99.999 pct, supplied by ALCOA. The metals were melted using a graphite or a sintered alumina crucible which was sealed in quartz tubes under an argon atmosphere. The alloys Were remelted to reduce segregation and homogenized in the 13 phase field for 24 hr. The weight of the alloys after homo-genization compared with the initial weight indicating the alloy composition to be within k0.1 pct of the in- tended composition. Chemical analysis of several samples confirmed the estimated alloy composition. All metallographic observations were performed on samples electropolished with ac current using a 10 pct solution of KCN and a stainless-steel electrode. The samples were cut into -1-mm-thick slices, polished, heat-treated in the ß phase field, and quenched in air, ice water, or iced brine. Some heat treatments consisted of a 2-min anneal carried out in air to allow maximum quenching speeds. Other samples were heat-treated in a vacuum of 10"4 torr and quenched using a helium blast in a quenching apparatus described elsewhere.&apos; In both cases, no appreciable losses of silver or aluminum were noted at the sample surface by sectioning and electropolishing. Table I summarizes the microstructure resulting from the various quenching treatments. The composition of the alloys investigated is also indicated on the equilibrium diagram in Fig. 1. The <, hcp phase, typical of a massive ß — cm, transformation, was observed in all alloys. A minimum temperature of 450°C was estimated for this transformation based on preliminary cooling experiments. the results being included in Fig. 1. In these

Jan 1, 1970

By A. M. Nicholas

WHEN attempting to lift mill pulp containing a considerable percentage of wolframite, in an ordinary bucket elevator, difficulty was encountered from the tendency of the tungsten minerals to settle, on account of their high specific gravity. Lifts not exceeding 20 ft. (6m.) could be accomplished without particular difficulty, because there was insufficient time for consolidation of the heavy minerals in the bottom of the bucket; but with higher lifts, the heavy tungsten minerals would settle so solidly that when the bucket turned over at the top of the elevator the contents could be only partially dumped. In this case it ,was necessary to lift the material a distance of 34 ft. (10m.); hence some means had to be adopted to prevent settling, at least to such an extent that the material could be dumped from the buckets. To reduce the carrying time to a sufficient extent would have required a belt speed of 475 ft. (144.7m.) per minute, which was not feasible. It was decided to try, by vibrating the belt about midway of its span, to keel) the material sufficiently agitated during elevation to prevent settling. Underneath the rising side of the belt was placed a 12-in. (30.48-cm.) idler pulley, 18 ft.-(5.5 in.) up in the total lift of 34 ft. The idler pulley was set eccentric by 3/4 in. (19 mm.) and it was hoped that the vibration thus imparted to the belt, running at a speed of 280 ft. (85 m.) per minute, would prevent settling in the buckets. It was found that the pressure of the belt on the idler, due to the small inclination of the elevator, was not sufficient to overcome sliding on the idler, and the agitation was insufficient to accomplish the desired purpose. The idler was then connected to the drive shaft of the elevator through an intermediate pulley, gearing down to a speed of 20 r.p.m., so that a positive vibration was imparted to the belt; this imparted enough motion to the belt to prevent settling and the contents of buckets clumped satisfactorily.

Jan 2, 1918

By R. Laird Auchmuty

FOREWORD THE A. I. M. E. Subcommittee on Bituminous Mining has been trying for several years to secure the information that was collected by the Marquette Cement Manufacturing Co. on the subsidence of its property caused by mining coal beneath it. The Marquette company was willing to have these data used, but it was impossible to obtain them because the records had been filed away. About two years ago these records became available and the engineering data presented herewith were compiled from them by R. L. Auchmuty, assisted by L. E. Young, who was connected with the collection of the information. On account of the cost, it is impossible to reproduce all of the data here, although the original information is on file with the records of the Committee on Ground Movement and Subsidence. The accompanying charts show representative and average data and give only the engineering information with enough description to tell how it was obtained. It is hoped that all of those connected with the original case will participate in the discussion that this will bring forth. HOWARD N. EAVENSON, Chairman, Subcommittee on Bituminous Coal Mining.

Jan 1, 1931

By J. E. Dutrizac, R. J. C. MacDonald, T. R. lngraham

When sintered disks of synthetic chalcopyrite (CuFeS2) were leached in acidified aqueous solutions of ferric sulfate, the following reaction stoichiometry was obtained: CuFeS2 + 2Fe2(SO4)3 = CuSO4 + 5FeSO4 + 2S Over the temperature range from 50º to 94ºC, the reaction displayed parabolic kinetics. The parabolic rate constant for the dissolution of copper is given by the equation: log.k(mg2/cm4-hr)= 11.850 - 3780/T The activation energy for the dissolution process is 17 ± 3 kcal per mole. The parabolic kinetics have been attributed to the progressive thickening of a sulfur film on the surface of the chalcopyrite. When the leaching solutions contain less than 0.01 molar Fe+3 , the Fe concentration influences the rate of leaching, probably through a mechanism involving the diffusion of ferric sulfate through the sulfur layer. At higher Fe+3 concentrations, the rate control in the leaching. reaction has been attributed to the diffusion of ferrous sulfate through the sulfur. The rate of reaction is insensitive to changes in acid concentration and in disk rotation speed. ThE reaction of acidic ferric sulfate solutions with various sulfide minerals is of practical interest for both bacterial and heap leaching. This leaching medium is generally used with low-grade ores that cannot be treated profitably by conventional means. In both bacterial leaching1-3 and heap leaching, the active agent for sulfide dissolution is ferric sulfate. Although the reactions of ferric sulfate with chalcocite, covellite, and bornite have been investigated,4*7 the kinetics of leaching chalcopyrite with ferric sulfate have not been thoroughly studied. This paper reports a study of that reaction. EXPERIMENTAL Reagent-grade sulfur was purified by the method of Bacon and FanelliB and then it was vacuum-distilled to remove any soluble magnesium salts that had been introduced during the purification procedure.9 From stoichiometric quantities of the purified sulfur and hydrogen-reduced electrolytic copper sheet (99.90 pct Cu), CuS was synthesized at 450°C in a vacuum-sealed, pyrex vessel. About 24 hr was required for the completion of the reaction. A similar procedure involving hydrogen-reduced iron wire (99.90 pct Fe) was used to synthesize FeS1.002. A 2-furnace arrangement was required. The iron was heated to 800°C while the sulfur was maintained at about 400°C. Although the reaction to consume the sulfur was rapid, the material required additional heating (1 week) in a sealed silica ampoule at 800°C before it was homogenized. X-ray powder diffraction analysis confirmed that the copper sulfide was covellite and that the iron sulfide was troilite. The composition of the iron mineral was confirmed by wet chemical analysis. The two sulfides were ground to minus 100 mesh, weighed in equimolar amounts, mixed thoroughly, and pressed into pellets at 80,000 psi. The pellets were vacuum-sealed in pyrex ampoules and then sintered for 3 days at 550°C after an initial heating at 450°C for a few hours. The pellets were then cooled, polished with 3/0 emery paper, rinsed in acetone, and stored. The material had the characteristic brassy color of chalcopyrite and was shown by X-ray diffraction to be CuFeS2. Microscopic examination of the polished surfaces revealed small inclusions of pyrite (approximately 0.5 vol pct) as the only impurity. The presence of small amounts of a second iron compound will not alter the amount of dissolved copper but might increase the amount of ferrous ion slightly. It was calculated that dissolution of all of the pyrite and 100 mg of Cu (a typical value) would change the expected ferrous concentration by only 4 pct. Microscopic examination of a pellet after leaching revealed that the pyrite was not preferentially solubilized; only those pyrite grains at the surface were attacked. Hence, the pyrite is unlikely to alter the rate of copper dissolution. The chalcopyrite disks were about 1.7 mm thick and 27 mm in diam. They were about 80 pct of theoretical density, and for this reason their true reaction area was somewhat larger than the 5.8 sq cm area presented by the polished face. The disks were cemented to lucite cylinders in such a way that only the polished face was exposed. The disks were then leached by methods previously described.6,7 RESULTS AND DISCUSSION Stoichiometry and Kinetics. The initial experiments were directed to the problem of resolving the stoichiometry of the leaching reaction. Disks of CuFeS2 were leached at 80°C for various periods of time in acidified ferric sulfate solutions that were protected from oxidation by a cover of flowing nitrogen. When the disks had been partly leached, they were removed, their soluble salts were washed out, and then they were treated with CS2 in a Soxhlet extraction apparatus. The ratio of elemental sulfur to dissolved copper thus obtained was approximately 2 to 1. After the extraction of elemental sulfur from the pellet, the residue consisted of unreacted chalcopyrite only. For runs in which an appreciable amount of copper was dissolved, the ratio of ferrous ion to cupric ion in the solution was

Jan 1, 1970