Search Documents

By Robert D. Thomson

Listed among strategic and critical minerals for stockpiling, mica is an important raw material in electronic and electrical equipment industries. Each year large quantities are consumed in the manufacture of television and radio tubes, radar sets, generators, transformers, electric toasters, and irons. Crude mica, which comes from the mine in crystals or in rough books of various shapes and sizes, is known as run-of-mine or book mica. Structural imperfections, impurities, rough edges, and defective laminae are removed by cobbing, rifting, and trimming to produce a marketable block mica or film mica. After the rough books have been hand-cobbed to remove adhering rock, dirt, and crushed or defective mica they are rifted, or split. In the U.S. a straight-bladed knife is generally used for rifting. Products of this operation are classed broadly as untrimmed sheet mica and scrap.

May 1, 1956

By Carl Zapffe

THE manganese ore reserves of Lake Superior, because of their location and nature, have recently achieved a marked degree of importance as compared with the world's manganese reserves. To appreciate fully the significance and the adequacy of this reserve, one should understand how such ores are used, where obtainable, the quantity used, possible avenues for increasing use or application, and the chemical and physical characteristics of the available ores. Only with that data at hand, is it possible to arrive at a convincing answer to the question whether the reserves of manganiferous iron ores are adequate to satisfy the annual demand of the presently known iron ores for the full life of the latter. That they should last is vital; and to ascertain whether they will last was the purpose of the present investigation, which marks the first attempt to study these Lakes Superior reserves. The Lake Superior region possesses and supplies the full assortment of iron ores needed for the conversion of ore into pig iron and steel ingots and castings. These ores are named bessemer; low-phosphorus and high-phosphorus non-bessemer, manganiferous and siliceous; and to include the term aluminiferous, would be in keeping with the most recent no men clature adopted by some dealers. About 35 or 40 years ago, perhaps less, most of the Lake Superior ore which is being used today was the material then considered either useless or too poor for immediate use; perhaps it was considered no more than a future reserve. An examination of the analyses of ores and the guarantees in the sales of those older times, convinces one that the opinions of those days must have been governed by only small differences in analyses. Some grades of ores used today in substantial amounts were used most sparingly only one or two decades ago. The relative quantities of the different ores used is ever changing; and, bit by bit, the metallurgy and the opinions are every changing and the prejudices against certain ores and practices decrease. This is unquestionably more pronounced for the manganiferous ores than for any other kind. If, when using the present data for manganiferous ores, we will only be

Jan 5, 1927

By W. S. Ayres

THE boiler from which these stay-bolts have just been obtained was that of the locomotive Catasauqua, Lehigh Valley Railroad, built at the company's shops, South Easton, Pa., in 1864. The iron is Lowmoor, and has been in use nearly ten years. The stay-bolts have become highly crystallized from some cause, and are very brittle, only requiring three blows of an ordinary machinist's hammer to break them. On the other hand, when annealed they are remarkably tough, as shown by the doubled sample. The majority of the broken bolts were found to be along the curve of the shell from a to c, Fig. 6. All of them were broken just inside the shell, and in no case next the fire-box. The fractures seen to have been of slow procedure, beginning at the top of the bolt, gradually working deeper and deeper until the pressure of steam overcame the tenacity of the remaining parts. Unequal contraction and expansion of the parts of the boiler seems in this case to be the only existing cause for a fracture. We cannot, from the data obtained, prove to an absolute certainty the manner in which the combined forces generated by the unequal contraction and expansion of the parts have arrived at their resultant work ; but we can, from these data and from existing circumstances, arrive at what seems to be a fair supposition, and thus at least anticipate the truth. The contraction and expansion in locomotive boilers has a large range between its extremes, and is very variable, both with reference to time and to their component parts. The shell and fire-box do not, perhaps, at any time, contract and expand at the same rate, nor proportionally ; but, on the contrary, owing to the great changes in temperature that take place both in the atmosphere around the shell, and in the fire in the fire-box, their condition is often changed suddenly, and widely from that which they just previously had. This racking tendency produces strains in the different parts, and it is fair to suppose that the continual motion of the two

Jan 1, 1874

By W. S. Ayres

The boiler from which these stay-bolts have just been obtained was that of the locomotive Catasauqtia, Lehigh Valley Railroad, built at the company's shops, South Easton, Pa., in 1864. The iron is Lowmoor, and has been in use nearly ten years. The stay-bolts have become highly crystallized from some cause, and are very brittle, only requiring three blows of an ordinary machinist's hammer to break them. On the other hand, when annealed they are remarkably tough, as shown by the doubled sample. The majority of the broken bolts were found to be along the curve of the shell from a to c, Fig. 6. All of them were broken just inside the shell, and in no case next the fire-box. The fractures seem to have been of slow procedure, beginning at the top of the bolt, gradually working deeper and deeper until the pressure of steam overcame the tenacity of the remaining parts. Unequal contraction and expansion of the parts of the boiler seems in this case to be the only existing cause for a fracture. We cannot, from the data obtained, prove to an absolute certainty the manner in which the combined forces generated by the unequal contraction and expansion of the parts have arrived at their resultant work ; but we can, from these data and from existing circumstances, arrive at what seems to be a fair supposition, and thus at least anticipate the truth. The contraction and expansion in locomotive boilers has a large range between its extremes, and is very variable, both with reference to time and to their component parts. The shell and fire-box do not, perhaps, at any time, contract and expand at the same rate, nor proportionally; but, on the contrary, owing to the great changes in temperature that take place both in the atmosphere around the shell, and in the fire in the fire-box, their condition is often changed suddenly, and widely from that which they just previously had. This racking tendency produces strains in the different parts, and it is fair to suppose that the continual motion of the two

By H. Shaw

Gravitational measurements made by means of the Eotvos torsion balance over any area enable a representation to be obtained of the total gravitational effects over the surface of that area arising from the combination of all sources of gravitational disturbance. The investigation of gravitational anomalies occurring in the upper layers of the earth's crust, in this way, has been attended with considerable success, and, especially in areas in which no undesirable disturbing factors were present, it has been possible to locate geological structures and mineral deposits with remarkable accuracy and also to delineate their confines and to indicate their depth below the surface with a certain amount of precision. Before this can be done successfully, however, it is necessary to determine the gravitational effect arising from all other extraneous anomalies in order that these may be satisfactorily eliminated. Disturbing features giving rise to these gravitational anomalies may be divided into two classes, according to whether they are situated at or below the surface. The former are determined readily by a detailed leveling of the ground surface, and may be satisfactorily eliminated provided the specific gravity of the terrain material is known, but the latter are generally unknown, and can only be ascertained with difficulty. When, therefore, the observations obtained by means of the Eotvos torsion balance have been corrected for all surface irregularities, known as terrain and topographical effects, there remain residual gravitational values which are due entirely to subterranean conditions. The. task of interpreting these results consists in disentangling the various subsurface anomalies, and subsequently in computing and recombining the effects of these various masses in order to verify the interpretation. When the gravitational anomalies are due to large and comparatively simple structures it is a matter of little difficulty to arrive at a satisfactory interpretation of the subterranean conditions, but in more complex areas, where the geological conditions are more complicated and several irregular features exist, the process of disentangling the various constituents of the complex mass arrangement is more difficult and requires the expenditure of considerable time and labor. When the subterranean effects have been obtained, they are subjected to a critical examination, and an effort is made to recognize and

Jan 1, 1932

By B. Post, D. Moskowitz, F. W. Glaser

Lattice parameters of various hafnium compounds, two borides, a carbide and a nitride were determined. Some physical properties were measured and comparisons with the properties of the corre-sponing zirconium compounds were drawn. A LTHOUGH borides of transition metals have A been extensively investigated in recent years, there are few references in the literature to borides of hafnium. This is not surprising, in view of the extreme scarcity and great cost of the purified metal. Becker' reported the existence of a boride, carbide, and nitride of hafnium; he listed the formulas as HfB, HfC, and HfN. However, in his work no mention was made of chemical analyses of the reaction products. In this paper results of a study of various hafnium compounds, including borides, carbides, and nitrides, are reported. It was expected, of course, that these compounds would closely resemble their zirconium analogues, which have previously been extensively investigated. Three borides of zirconium are known: ZrB,² ZrB,² and ZrB212.4 In addition, one carbide, ZrC,5 and one nitride, ZrN,5 have been reported. It should be noted that ZrB and ZrB12 are stable phases only in relatively high-temperature regions, unlike ZrB2, which is stable from room temperature to its melting point." It was felt that hafnium would yield corresponding borides, carbides, and nitrides with similar stability ranges. Experimental Procedures and Results A quantity of hafnium metal that was better than 97 pct pure, with zirconium as its major impurity, was furnished by the Argonne National Laboratories. The boron powder used was approximately 96.5 pct pure. In general, the procedure involved mixing the metal and boron powders in the desired proportions, and subsequently hot pressing them. The resulting specimens were analyzed chemically and phases present were identified by X-ray diffraction methods. A Philips Norelco Geiger counter spectrometer was used for identification of components and for measurement of lattice constants. Filtered copper radiation was used for all X-ray work. Electrical resistivity and density measurements were carried out at room temperature. The stable diboride, HfB2, will be described first. Stoichiometric quantities of boron and metal were heated to approximately 2000°C during hot pressing. The reaction product was shown by X-ray diffraction methods to be isomorphous with ZrB2 (space group Dlgh; P6/mm). Values of lattice constants of HfB2 compared with those of ZrB2 are shown in Table I. The calculated density of HfB2 is 11.2 g per cc, compared with a measured (pycnometric) value of 10.5 g per cc. HfB2 specimens prepared under a wide variety of conditions (including variations in initial Hf-B ratios) showed only insignificant variations in lattice constants, indicating a very narrow range of homogeneity. Like ZrB2,7 HfB2 is an excellent electrical conductor. The resistivity at room temperature of hot-pressed samples, approximately 85 pct dense, is 12 microhm-cm. This compares favorably with the electrical resistivity of 10 mi-

Jan 1, 1954

By D. W. Mitchell

J. H. Schlobohm (Joy Manufacturing Co., New York) —This paper has been read with a good deal of interest; however, there are several salient features which Mr. Mitchell has overlooked. The initial emphasis on selection should be placed on type of material to be loaded and method of mining involved. There is no possible connection in the comparison of scraper or Joy Lohite loaders with loaders of any other type. Neither can an operating comparison be made between "Rocker type" loaders and those equipped with loading head and gathering arms. Each unit has its application, but the application depends on material and the mining system involved. Mr. Mitchell's compilation of data on various loaders, while it represents a great many facts, does not represent true operational data, i.e. Table I, column heading "Max. Rated Capacity for 7 hr;" these figures are meaningless inasmuch as little or no time at all is allowed for tramming or switching or moving the loader from place to place. Table I, column heading, "Best Reported Field Shift," would be a better source of operational data, assuming of course that conditions, applicable to each class of loaders, were average to avoid enthusiastic performance figures. Table I, column heading, "Repair Parts Cost per Ton," is still another item which will not stand up under comparison unless the loaders involved are put into type classes, i.e., there are simply no means possible to show the repair costs of scraper loading versus a caterpillar mounted loader in abrasive material. The source of information here is also important. Actual field tests have proved that these costs are of a constant nature. Manufacturers are striving to hold them to a minimum and many interesting studies have been made. A fairly recent example of this in a large United States gold producer has proved that Joy HL-3 loaders can be consistently average, less than $0.025 per ton mucked. This cost represents repair labor and parts over a two year

Jan 1, 1952

By Charles R. Keyes

PAGE. I Introduction,........674 11. Occurrence............ 676 III. Geology of Death Valley BoRate-Region,....677 1. Surface.Relief ......677 2. Geologic Formations,....680 3. Geotectonics,..684 IV. FURnace Canyon Deposits..686 1. General Features..686 2. Geologic Structure..688 3. Ores.....690 4. Typical Section ...691 V. Lost Valley Deposits...693 VI. Mojave Desert Deposits..693 1. Distribution....693 2. Geologic Structure ........694 3. Ores............ 696 VII. Santa Clara Valley Deposits .699 1. Distribution....699 2. Geologic Structure..700 3. Ores.....700 VIII. Ventura County Deposits..701 IX. Geologic Occurrence of Borax .701 1. Original Sources...701 2. Solfataric Borax...702 3. Lacustrine Borax...703 4. Marsh Borax....703 5. Terranal Borax...704 X. Chemistry of Natural Borates.... 705 1. General Considerations.....705 2. Composition of Saline Waters...706 3. Solubility of Components....707 4. Time-Element in Water.Concentrations . 707 5. Effect of Temperature.....708 6. Effect of Pressure......709 XI. Occurrence of Other Commercial Salines,.... 709 I. Introduction. A complete transforrnation has taken place in the borax industry during the year 1908. A most remarkable factor in

Jan 1, 1910

Members whose deaths were reported from Mar. 5, 1919, to Apr. 5, 1920. Elected Died 1895 ANDERSON, ROBERT HAY 1920 1916 ATWATER, M. W 1919 1905 BARD, D. C 1920 1893 BECKER, GEORGE F 1919 1914 BIGELOW, BRAXTON 1917 1917 BLACK, HERBERT F 1919 1911 BOWLER, ROBERT P 1919 1902 BROWN, ARTHUR H 1919 1913 BURNHAM, WILLIAM 1918 1905 Carnegie, Andrew 1919 1900 CARPENTER ROLLA C 1919 1908 CONNER H. MCKEAN 1919 1883 COOK, ROBERT A 1919 1903 COOPER, JOHN 1919 1875 CROXTON, SAMUEL W 1919 1889 DICKMAN, R. N 1919 1893 DUFOURCQ, EDWARD L 1919 1917 FAY, CHARLES H 1918 1875 FOHR, FRANZ 1919 1879 FRICK, H. C 1919 1880 GAYLEY, JAMES 1920 1917 GARVIN, JOHN C 1920 1915 GIBBS, CHARLES H 1919 1917 GODBE, RALPH T - 1918 GREGORY, A. E 1919 1917 HAMILTON, A. McL 1918 1914 HEISE, HENRY C - 1913 HINCKLEY, E. R 1918 1884 JENNINGS, HENNEN 1920 1895 JOHNSON, J. E., JR 1919 1895 KANN, MEYER M 1919 1895 KAYSER, H. W. F 1919 1918 KENNEDY, DANIEL J 1903 KING, PAUL S 1919 1901 LANGTON, JOHN 1920

Jan 1, 1923

By C. E. Morris, S. R. Pollack

Current -vo1tage -temperature characteristics were studied for Al-A12O3-Al, Au structures. The oxide film was grown by gaseous oxidation in an 0, glow discharge. The electron-transfer mechanism was identified as tunneling and the barrier heights and thickness were determined. It was found that the barrier height at the parent aluminum-oxide interface is 1.5 ev and the built-in vloltage with an aluminum counter electrode is 0.35 v. The growth mechanism of the oxide is also discussed and it is suggested that Mott&apos;s theory is not applicable in describing the growth. Discrepancies were observed in the detailed comparison of tunneling theory, based on a trapezoidal model, with the data, and several possible explanations are discussed. THEORETICAL analyses of electron tunneling from one nonsuperconducting metal to another through an extremely thin insulating film have been presented.1-6 These calculations relate the current density, J, to the voltage, V, between the two metals, and to the temperature, T. The results of these calculations indicate that the magnitude of J and the detailed shape of the J-V characteristic will in general depend most strongly upon such properties of the insulating film as the thickness, the dielectric constant, and the structure of the electron energy barrier profile. The experimental results of many7-11 workers have corroborated the main features of these calcu- lations and have in fact been used to determine properties of the thin insulating film. It has been demonstrated, for example, that the tunneling current formulation of simmons4 is adequate to describe the observed J-V characteristic over approximately 10 orders of magnitude of current. Furthermore, from these observations and analyses, insight into the nature of the A1-A12O3 (thermally grown) interface region has been obtained.&apos;&apos; The temperature dependence of the tunnel current has also been studied and, although a self-consistent set of tunneling parameters can be chosen to describe these data at low voltages (i.e., <0.5 v), the tunneling parameters are different from those obtained from higher voltage (i.e., of the order of 1 to 5 v) measurements." We have measured the effect of temperature and voltage upon the tunnel current at higher applied voltages and find that the theory is inadequate to describe all of the effects reported here. A discussion of the origin of this inadequacy is presented following a description of the samples used for these experiments, and the experimental results. SAMPLE PREPARATION Metal-insulator-metal structures were fabricated in the following way. A 99.999 pct A1 charge was used to deposit, in vacuo, thirty films, 2000Å thick, onto a 1 by 2 in. glass substrate (7059 Corning glass). Prior to deposition the substrate was glow-discharged for final cleanup and then baked at 200°C for 2 hr at a pressure of 5 x 10-7 torr. The pressure during deposition of the metal films was 2 X 10-6 torr and the deposition rate was 50 to 100Å per

Jan 1, 1965

By H. L. Brunjes, G. B. Gould

The softening temperature of coal ash, as determined in the laboratory, has been used for years as an indication of the tendency of coal to form clinker and slag. It has not, however, provided an index of unfailing and complete reliability for forecasting the relative performance of different coals, even in the same plant and under the same conditions of operation. A large amount of investigation has been done in attempts to account for this, but practically all of this has been in the direction of studying the chemical constitution of coal ash. Much has been learned, but without much progress toward a better practical application of knowledge to the selection of coal. There has always seemed to be an important factor that eluded discovery. It has long been recognized that the ash-softening temperature as determined in the laboratory had one weakness, in that it represented the temperature at which a thorough mixture of the whole ash of the coal would melt, the mixing having taken place before the heal was applied. This does not, of course, simulate what takes place in practice. With certain exceptions, which will be noted later, almost all of the investigations that have been made, especially the chemical studies of coal ash, have likewise dealt with a mixture of the whole ash in the coal. The present investigation differs from these fundamentally in its approach to the problem. It started from the observation that some clinkering and slagging obviously takes place in a boiler furnace at temperatures below the laboratory-determined ash-softening temperature of the coal. Knowing that coal as it is shipped to market is a heterogeneous mixture of diverse mineral substances, each of which may act independently of the others in the initial stages of clinker and slag formation, it seemed logical to investigate separately the softening temperature of the ash in physically separable portions of the coal, and to determine the relative quantities of ash contributed by each portion. While the data are far from complete, covering as they do only a few coals, and while the conclusions to be reached from the data at this stage arc, in some measure, speculative, they are presented at this time because

Jan 1, 1940

By G. B. Gould, H. L. Brunjes

The softening temperature of coal ash, as determined in the laboratory, has been used for years as an indication of the tendency of coal to form clinker and slag. It has not, however, provided an index of unfailing and complete reliability for forecasting the relative performance of different coals, even in the same plant and under the same conditions of operation. A large amount of investigation has been done in attempts to account for this, but practically all of this has been in the direction of studying the chemical constitution of coal ash. Much has been learned, but without much progress toward a better practical application of knowledge to the selection of coal. There has always seemed to be an important factor that eluded discovery. It has long been recognized that the ash-softening temperature as determined in the laboratory had one weakness, in that it represented the temperature at which a thorough mixture of the whole ash of the coal would melt, the mixing having taken place before the heal was applied. This does not, of course, simulate what takes place in practice. With certain exceptions, which will be noted later, almost all of the investigations that have been made, especially the chemical studies of coal ash, have likewise dealt with a mixture of the whole ash in the coal. The present investigation differs from these fundamentally in its approach to the problem. It started from the observation that some clinkering and slagging obviously takes place in a boiler furnace at temperatures below the laboratory-determined ash-softening temperature of the coal. Knowing that coal as it is shipped to market is a heterogeneous mixture of diverse mineral substances, each of which may act independently of the others in the initial stages of clinker and slag formation, it seemed logical to investigate separately the softening temperature of the ash in physically separable portions of the coal, and to determine the relative quantities of ash contributed by each portion. While the data are far from complete, covering as they do only a few coals, and while the conclusions to be reached from the data at this stage arc, in some measure, speculative, they are presented at this time because

Jan 1, 1940

By A. A. Collins

When we speak of high copper on a lead blast furnace we think in terms of 4 to 5 pct, or. any lead charge carrying over 1 pct. Any copper on charge will produce its corresponding troubles such as lead well, extra slag losses, drossing problems, and the working up of the dross. This is indeed a very interesting subject and one that used to give the old-time lead metallurgists such as Eiler, Hahn and lles many worries, not so much in the actual operation of the hlast furnace but in the working up of the copper. When the American nletallurgists commenced with the American rectangular-shaped lead blast furnace in the 1870&apos;s and got away from the reverberatories such as were in use in Germany and other parts of the world, they went to greater tonnages, as 80 to 100 tons per day in comparison to the 20 to 30 tons per day in the other processes. With the greater tonnages along with insuficient settling capacity, the silver losses in some cases were increased. Hence the lead-fall was low, for there were no leady concentrates in those days to assist the metallurgist to gain lead or an absorber for the precious metals; and in some cases copper sulphides were added intentionally to the charge to produce a copper matte to lessen the silver losses through the dump slag. The operators in those days thought that where some copper was always present in the lead ores the copper should not enter into the reduced lead and alloy with it. This, by the way, is just the reverse of our present-day practice, when we try to put all of the copper into the blast furnace lead and to remove the same through the drossing kettles. Therefore the furnace was operated to produce a certain amount of matte or artificial sulphides, since, due to the great affinity of copper for sulphur, any copper present would enter the matte almost completely. Thus, the lead bullion produced was practically free from copper. The products of the furnace were metallic lead or lead bullion, containing 05 to 95 pct of the lead and about 96 pct of the silver which were in the ore—a lead-copper-iron matte which contained nearly all the copper in the ore and the slag, the waste product. In the United States, up through the year 1092, we find the small furnace 100 X 32 1/2 in. with 12 tuyeres, some 6 on each side, plagued with a small amount of poorly roasted sulphides— either from heap or hand roasters that produced matte. This matte was roasted and if poor in copper was returned for the ore smelting. Otherwise it was smelted either alone or with additions of rich slags or argentiferous copper ores, the products being lead and a highly cupriferous matte, the latter being subsequently worked up for its copper. The lead metallurgists kept trying and improving on furnace and roasting equipment designs until we find blalvin W. Iles constructing at the old Globe Plant at Denver what came to be the modern furnace. That is, in 1900 he built a furnace of 42 in. width by 140 in. at the tuyeres with a 10 in. bosh and a 16-ft ore column. This type has been more or less standard to the present time, though modified in width and length to meet the demand for large tonnages and improvements in structural details. In 1905 at Cananea, Mexico, Dwight and Lloyd developed the present down-draft sinter machine that has meant so much in producing a well-processed material for the lead blast furnace. In 1912 Guy C. Riddell came forth with double roasting at the East Helena Plant of the American Smelting and Refining Co., which removed the "zinc mush plague." Incidentally, with the introduction of double roasting, which most lead plants were forced into after 1924, when lead flotation came into its own, less matte or no matte was produced. When this stage arrived, the copper was forced into the dross and the casting of lead at the blast furnace lead-wells was stopped. In plants with a fair copper carry 1 pct or better on the blast furnace charge, the lead wells became inoperative once the production of matte stopped. The copper drosses clogged the lead wells and even with bombing, either water or dynamite, the operators could not keep them open. Thus, the lead wells were abandoned in some plants, such as at the El Paso and Chihuahua smelters of the American Smelting and Refinillg Co., and all lead taken out through the first settlers. The elimination of sulphur, espccially sulphide sulphur, from the blast furnace charge and the nonproductiori of matte resulted in a great saving of tinie, energy and equipment in the recirculation of the copper, With the copper content in the dross and dross-fall ranging in quantities from a few percent up to 60 pct, such as at El Paso, a drossing problem was created. As the old-time operators hated dross and buried the same in the shipping bullion, the modern metallurgists from 1925 on decided that with increasing dross-falls they would have to adopt the lead refiner&apos;s ideas of drossing kettles with subsequent treatment of the lead with a sulphur addition to have the shipping lead of 0.01

Jan 1, 1950

By F. Habashi

Conclusive evidence is given showing that the dissolution of gold in cyanide solution follows mainly the over-all equation A similar equation can also be written for the dissolution of silver. Theoretical derivation of the velocity equation for this reaction has been obtained, which describes quantitatively the experimental facts. The equation is as follows: where [CN-] and [O2] = the concentrations (in moles/ml) of cyanide and dissolved oxygen, respectively. DcN- and Do2 = the diffusion coefficients of cyanide and dissolved oxygen; 1.83 x 10-5 and 2.76 x 10-5 cm2 sec-1, respectively. A = the surface area of the metal in contact with the aqueous phase, in cm S = the thickness of the boundary layer; varies between 2 and 9 x 10-3 cm, depending on the speed and method of agitation. Rate being expressed in g. equiv. sec-1. INTRODUCTION During recent years a number of papers have been published on the kinetics of dissolution of gold and silver in cyanide solution.1,6 The authors agree that the rate of dissolution is diffusion-controlled, but disagree regarding the mechanism of the reaction. Some authors&apos;-3 support the Bodlander equations: The apparently conflicting results, when critically reviewed, are concordant and lead to the conclusion that the proper equation for the dissolution reaction should be only Eq. 1. This conclusion is based on the following facts: 1) For every 2.2 equivalents of metal dissolved one mole O2 was consumed. 2) For every 1 equivalent of metal dissolved, 2 moles of cyanide were consumed.&apos; 3) Hydrogen peroxide is formed during the dissolution of gold or silver, and for every 2.2 equivalents of metal dissolved, one mole H2O2 was produced.1,2 The fact that Kudryk and Kellogg4 did not detect H2O2 during the dissolution of gold should not throw doubt on the many affirmative reports of other investigators. Lund2 has shown that depending on the crystalline nature of the silver surface, (e.g. method of etching the sample) H2O2 may undergo catalytic decomposition heterogeneously according to: Further Kudryk and Kellogg4 carried out their experiments in the presence of 0.5% KC1 to achieve a greater electrolytic conductivity, and Matsudaira7 has shown that H2O2 is catalytically decomposed (homogeneously) in the presence of C1- ion in alkaline solution. 4) Experiments carried out by Boonstra8 and Lund2 showed that the dissolution of gold and silver respectively in KCN + H2O2 in the absence of oxygen was a slow process. Thus the reaction takes place to a minor extent only. Further kinetic studies showed that: 1) The rate of dissolution depends on the surface area of the metal in contact with the liquid phase, thus indicating that the process of dissolution is a heterogeneous process. 2) The rate of dissolution depends on the speed of stirring,5,6 thus indicating that the process is controlled by a physical phenomenon.

Jan 1, 1967

By Frederic E. Carter

THE gold, silver, copper alloys have been the subject of several fairly complete investigations by Jänecke, Sterner-Rainer1 and others, and indeed it would seem as if almost too much labor had been expended on them when one thinks of the numerous other ternary alloys calling for more light; however, the combinations of the three metals are actually so widely used in jewelry that they merit perhaps the work done on them. There are a few properties of importance and interest not treated by the authors mentioned which the writer investigated some little time ago, and therefore he has ventured to put on record the results of some of his tests made at that time. He has chosen, as being the two series of most importance to the jeweler, only the 14 and 18-kt. golds, so the subject is by no means so inclusive as that previously treated. As stated, this paper is merely a record of tests made in a practical manner on two useful series of alloys, and is not at all a critical study of their properties. 14-KARAT ALLOYS These alloys are important because they are the most generally used in 14-kt. jewelry; by changing the proportions of silver and copper there are obtained all shades of color from the green of the alloy containing gold and silver but no copper, to the red of the alloy containing gold and copper but no silver. Many 14-kt. gold alloys are made with the addition of zinc to the three metals mentioned here; the added metal does not appreciably alter the color of the alloy nor has it, in general, much effect on the working properties; it may, however, prove useful as part of the zinc added may act as a deoxidizer and render the alloy free of gas. It is not a necessary addition because with proper precautions all the 14-kt. gold alloys containing only silver and copper in addition to the gold can be made free of gas and capable of being cold rolled without cracking. The alloys in which approximately equal quantities of silver and copper

Jan 1, 1928

By Zárate G. E.

Practical and statistical considerations concerned with the sampling, sample preparation and chemical analysis of precious metal ores are reviewed. The principal requirements related with the characteristics of sample cutter, sample preparation, equipment and methods are discussed. The main errors introduced in the different stages of the sampling systems are defined. The application of statistical techniques to evaluate these errors to determine the ore grade variation and to design the number of increments needed to get an accurate assay are presented. Finally, results obtained with the application of this method to gold and silver ores are analyzed and some recommendations are given.

Jan 1, 1984

By Frederic Carter

THE gold, silver, copper alloys have been the subject of several fairly complete investigations by Jänecke, Sterner-Rainer1 and others, and indeed it would seem as if almost too much labor had been expended on them when one thinks of the numerous other ternary alloys calling for more light; however, the combinations of the three metals are actually so widely used in jewelry that they merit perhaps the work done on them. There are a few properties of importance and interest not treated by the authors mentioned which the writer investigated some little time ago, and therefore he has ventured to put on record the results of some of his tests made at that time. He has chosen, as being the two series of most importance to the jeweler, only the 14 and 18-kt. golds, so the subject is by no means so inclusive as that previously treated. As stated, this paper is merely a record of tests made in a practical manner on two useful series of alloys, and is not at all a critical study of their properties. 14-KARAT ALLOYS These alloys are important because they are the most generally used in 14-kt. jewelry; by changing the proportions of silver and copper there are obtained all shades of color from the green of the alloy containing gold and silver but no copper, to the red of the alloy containing gold and copper but no silver. Many 14-kt. gold alloys are made with the addition of zinc to the three metals mentioned here; the added metal does not appreciably alter the color of the alloy nor has it, in general, much effect on the working properties; it may, however, prove useful as part of the zinc added may act as a deoxidizer and render the alloy free of gas. It is not a necessary addition because with proper precautions all the 14-kt. gold alloys containing only silver and copper in addition to the gold can be made free of gas and capable of being cold rolled without cracking. The alloys in which approximately equal quantities of silver and-copper

Jan 1, 1928

By E. F. Steeb, P. C. Rosenthal

BlANKS 8 in. long were cut from the as-received bar stock and austenitized at 1560° F for 1 hr, oil quenched, tempered at 1140°F for 1 hr, and water quenched. The heat-treated blanks were machined into tensile specimens which were subsequently used for the straining operation. Some of the tensile specimens were temper embrittled at 900°F for 24 hr and air cooled prior to or following the straining operation. The specimens were strained in tension up to 5 pct total strain (elastic plus plastic) either at room temperature or at 900°F. The reduced sec- tion of the tensile specimen employed was 5 in. long and 0.600 in. in diam. This size specimen was sufficiently large to permit two standard Charpy impact bars to be machined from it. One group of specimens was temper embrittled at 900°F for 24 hr under stress of about 20,000 psi applied with a hydraulic tensile testing machine. Approximately 0.34 pct plastic strain occurred during this treatment. A comparison of series "A" and "B" data in Table I at 0 pct strain, reveals that the steel was rather highly susceptible to temper embrittlement, the embrittling treatment causing the transition temperature to rise 165°F. The 28 ft-lb level, which was near the center of the sharp-rising portion of the transition curve, was selected as the

Jan 1, 1959

By E. C. Houston

The phosphate ores mined in middle Tennessee typically consist of granular rock phosphate particles disseminated in a clayey matrix. In the TVA plant near Columbia, Tenn., the phosphate ore is mined, made into a slurry with the addition of a small amount of sodium hydroxide as dispersant, and treated in a hydroseparator to remove minus 10 micron material. The hydro-separator underflow, comprising a rough concentrate, is transported by pipeline to the plant; the hydro-separator overflow, comprising a tailing, is flocculated by addition of calcium sulphate and is discharged to settling ponds at a rate of about 1400 gpm. Sedimentation in the ponds produces a clarified effluent, which may be recycled for use as process water or discharged to surface drainage. This method of tailing treatment is not entirely satisfactory since poor sedimentation characteristics of the tailing result in poor ultimate utilization of pond storage volume.

Jan 1, 1949

By J. F. Shirley, H. K. Burke

In January, 1964, three years of laboratory and pilot plant test work reached its culmination. The San Manuel Div. of the Magma Copper Co. switched its molybdenite extraction circuit from the sodium hypochlorite-ferrocyanide process to the new process described in this paper. The San Manuel Div. Concentrator of the Magma Copper Co. is located near San Manuel, Ariz., approximately 45 miles northeast of Tucson. The concentrator treats approximately 35,000 tons of ore per day containing about 0.85% copper and 0.025% molybdenite. The San Manuel ore body is a disseminated copper deposit in quartz monzonite and quartz monzonite porphyry. The primary copper mineral is chalcopyrite; minor amounts of chalcocite, bornite and covellite are also present in varying degrees. Pyrite makes up somewhat more than half the sulfide content of the rock. The molybdenite is nearly always found on fracture faces or associated with quartz veinlets.&apos; On January 9, 1964, the sodium hypochlorite-ferrocyanide process that had been used for molybdenite recovery was replaced by the following process: 1) A conditioning step with hydrogen peroxide, sulfuric acid, sodium cyanide and zinc sulfate. 2) A two stage rougher flotation step with stove oil and sodium ferrocyanide. 3) One cleaning step with sodium ferrocyanide. 4) One cleaning step with sodium hypochlorite and potassium ferricyanide. 5) Four cleaning steps with an anti-foam Exfoam 636 and potassium ferricyanide. COPPER CONCENTRATOR Since molybdenite recovery is closely allied with previous steps in the operation, the copper concentrator will be briefly described first. The ore is ground to approximately 5% plus 65-mesh and 65% minus 200-mesh in a two-step grinding circuit consisting of rod mills in open circuit and ball mills in closed circuit with rake classifiers. A straight forward rougher circuit producing only one concentrate is used. The rougher concentrate is combined with the discharge from the regrind ball mills and is then cycloned, with the overflow going directly to the cleaner circuit and the underflow going to the regrind mills. The feed to the cleaner circuit is approximately 90% to 94% minus 325 mesh. The cleaner circuit consists of two stages in which the tailing from the final cleaner stage is returned to the primary cleaner and the primary cleaner tailing is returned to the rod mill discharge at the head of the circuit. Approximately 0.013 lb of the allyl ester of amyl xanthate (S-3302) per ton of ore is added to the rod mills and approximately 0.006 lb of sodium isopropyl xanthate (2-11) per ton of ore is stage added down the rougher banks. The frother used is methyl iso-butyl carbinol (MIBC) and it is added at various points in the grinding and the rougher flotation circuits. Lime is added to the ball mills and/or to the rod mills to maintain the desired pH in the rougher flotation. The rougher flotation has been operated at various pH levels between 10.5 and 11.5. Fuel oil is added in the regrind mills at a loading of approximately 0.01 lb per ton of ore. Factors Which Directly Affect Molybdenite Recovery: It has been found at San Manuel that, to obtain maximum molybdenite recovery, an oily type flotation collector must be added in the grinding circuit. When molybdenite is crystallized in the characteristic platy form, which appears to be the case almost all of the time at San Manuel, its surface has a greater tendency to adhere to oils than water. 2 Of the several oily type collectors available on the market, the allyl ester of amyl xanthate seems to be the best molybdenite promoter. The floatability of San Manuel&apos;s molybdenite does not appear to be affected, at least up to a pH of 11.5, by lime. It has been reported that at several plant operations high lime alkaline circuits have adversely affected the molybdenite recovery.3,4 It has been found through microscopic studies that the bulk of the chalcopyrite will float before the molybdenite in both the rougher and cleaner circuits. It is necessary to add fuel oil to the regrind mills to enhance the

Jan 1, 1965