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By Bunting S. Crocker

DISCUSSION Fred C. Bond (Allis-Chalmers Manufacturing Co., Milwaukee)—This is a very complete description of the screened ore fine grinding installation at Lake Shore. The only important data which might be desired are the operating screen analyses and power consumption in each stage. The statement that the energy consumption, or kw-hr per ton, remains constant for all mill diameters is of great interest, but the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases. More definite test work is needed to establish this point. Screened ore fine minding has been successful on hard Lake Shore ore &d on about equally hard South African Rand ore. However, it is still not determined whether it would be successful on ores of average or less than average hardness. It is possible that with softer ores the consumption of ore pebbles might be exceedingly high, or might even exceed the available supply. Also, the softer feed might result in a rapid buildup of harmful intermediate sizes in the mill circuit. These factors can be determined only by a full scale or long pilot mill test. In any case the Lake Shore staff is to be congratulated for this pioneering work. Bunting. S. Crocker (author's reply)—As requested, the operating screen analyses and Bower consumption for each of the four stages of pebble milling are given, see Table XII. The fist three stages are in series circuit; the last stage is in closed circuit. The primary circuit had rock feed averaging 124 g per piece and the final stages rock feed weighing 63 g. The intermediate stages had mixtures of the two. Regarding Mr. Bond's statement that "the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases," I would say that I have frequently been told this but have never yet been shown any evidence or seen any evidence published that would prove this statement. It may be true that some of the big mills recently installed have mechanical power savers such as micarta bearings and better lubrication, but as far as the power consumed inside the mill is concerned, we at Lake Shore think that it varies as D2 .", and that is all. We published extensive information on this point in our 1939 paper. Recently Mr. Banks at Consolidated published the figures on his 10 ft 8 in. x 12 ft 0 in. rod mill. As reported in this discussion by the writer: we calculated the power necessary to run this mill based on our 7x6-ft rod mill to within 1 pct of the actual power reported by Mr.

Jan 1, 1953

By T. J. Bosworth

The effect of hydrogen pretreatment of alumina, magnesia, and zirconia crucibles on reduced oxygen levels in iron charges has been studied. Hydrogen treatment of commercial high-purity magnesia crucibles was observed to produce as high as 70 pct reduction in oxygen content in iron melts. Hydrogen treatment of the commercial high-purity grade zirconia crucibles produced a 25 pct reduction in oxygen levels in iron melts. An average reduction of Knoop hardness of 47 pct was recorded on the iron charges melted in hydrogen-treated crucibles. On the basis of the experiments and the free-energy diagram, it appears that the major source of reducible oxygen to the iron charge is the suboxides and impurity oxides rather than the major oxide of the crucible. Suitable pretreatment of the crucibles to reduce and remove these oxides yields low oxygen levels in iron melted in the crucibles. DURING the last decade, attempts to produce more pure and uniform alloyed metals have brought about the development of many new melting techniques. Notable among these are vacuum-induction melting, inert and consumable-electrode melting, the cold-mold technique, and numerous techniques for removal of impurities by chemical reaction. However great the achievements in this field, metals must still be melted in contact with media of varying reactivity; i.e., gas, ceramic, and slag. Studies have been made to attain an understanding of the reactions that take place during melting, including slag-metal reactions and equilibria, equilibrium of hydrogen and water vapor, and carbon dioxide and carbon monoxide with oxygen in molten iron. These reactions have been treated quite thoroughly. There is one condition, however, that has been widely overlooked, and, to the author's knowledge, only cursive investigation has been made of it. This is the reaction between the ceramic crucible material and the iron charge. In recent years, Hadley and Bianchi' made a series of studies on hydrogen deoxidation of iron in magnesia crucibles. Bianchi2 studied crucible-FeO equilibria in the temperature range 1525°to 1650°C in high vacuum with, however, inconsistent results. Richardson and Jeffes3 in their series of articles on thermodynamic reactions in the making of iron and steel refer to Wentrup and Hieber's4 work on iron and aluminum deoxidation. Their pork included melting iron in alumina crucibles and measuring the aluminum and oxygen content in the iron. Because of the low values obtained, question has arisen of their analytical accuracy.3 Dastur and chipman5 obtained values on hydrogen-oxygen equilibrium in molten iron contained in alumina crucibles similar to those obtained in experiments in the present program. Hochmann6 studied removal of impurities in chromium irons in high vacuum. His results with respect to oxygen level in chromium-iron melted in magnesia crucibles are consistent with results we have obtained.

Jan 1, 1959

By J. Szekely

A rrlethod is presented for the estitnation of oxygen pickup by teettzed molten steel streams. Of the mechanisrt~s considered for oxygen absorption, physical entrainment appeared to be the most significant. The estent of this physical entrainment may be calculated by the application of boundary layer theory. It was found that the values of oxygen pickup thus predicted were in wry good agreement with data reported in the literature, in the case of short, well-defined metal streams. This entrainment model does not provide an adequate description of oxygen pickup by long streams, exoosed to external winds and where breakup of the streatn is likely. WheN molten steel is teemed into a mold or tundish it will absorb oxygen contained in its environment. Recently interest has been expressed in the measurement and possible prediction of the extent of this oxygen absorption: this interest owes its existence to two principal factors: a) the desirability of providing a quantitative assessment of the role played by the absorbed oxygen in impairing the quality of (vacuum degassed. stainless, or continuously cast) steel; b) the need for providing guidelines for stream protection methods. The purpose of this communication is to suggest a simple method for the prediction of oxygen absorption by continuous metal streams and to compare the predicted values with the data available in the literature. It is hoped that this method may provide at least an initial framework for the interpretation of additional data that might become available. FORMULATION Consider a continuous metal stream falling into a molten metal pool as depicted in Fig. 1. This system may absorb oxygen due to the following two mechanisms: 1) physical entrainment of gases (including oxygen) into the molten pool where reactions. thus absorption, will occur; 2) mass transfer (convec:tive diffusion) of oxygen to the metal stream. 1) Physical Entrainment. The physical entrainment may be calculated by using the concept of the diaplace-ttzent tlzickness of the laminar boundary layer, which will be stated briefly in the following.&apos; When a fluid flows past a solid surface, a portion of the fluid close to the surface is retarded due to fric-tional forces. This region in which the velocity of the fluid differs from that in the main stream is called the boundary layer. An illustration of this is given in Fig. 2 where it is seen that the thickness of the boundary layer increases as we move along in the x direction. In the laminar region, when Re, < 10&apos; the thickness of the boundary layer is given by the following relationship: f =4.64 Re;1" [1] where 61 is the boundary layer thickness, x is the length along the leading edge, Us is the velocity of the free stream, (i is the viscosity of the fluid (gas in this instance).

Jan 1, 1970

"The richest hill on earth" had its start as a placer gold camp in 1864. The gold was quickly worked out but then silver was discovered in the copper ores in Butte Hill. After ten years as a silver camp, rich copper ore was found in 1881 and the future of Butte was assured. In the early 1960&apos;s, the Anaconda Company began the shift from underground mining at Butte to open pit mining of the enormous, low-grade copper deposit surrounding the old high-grade veins which had been worked for over eighty years. Their first primary crusher was a 66 in. by 84 in. (168 m by 2. 13 m) jaw crusher. The pit was expanded at such a rate, however, that its location became endangered. Additionally, the new Clyde Weed concentrator had been built at Butte; and in 1969 construction of a new gyratory installation was started. The primary crusher at the Clyde Weed concentrator is one of the most unusual primary crushers ever put into operation. At times, the ore from the Berkeley pit contains large amounts of timber, pipe, rails, and so forth left in the ore body from the old underground mining operations; and the crusher installation had to be designed so this tramp material could be readily removed ahead of the crusher. Ore is delivered by 100- and 150-ton trucks to two side-by-side dump pockets, each with a capacity of 400 tons. Originally, a hydraulic grapple was installed on the dividing wall between the dump pockets; but it was found that oversize boulders large enough to block the discharge opening at the bottom of each hopper were rapidly covered with other material, and the grapple was unable to reach and remove the large boulders causing the blockage. In addition, the grapple was located in a quite vulnerable position; and it sustained damage from being struck by rock dumped from the trucks. Oversize boulders are now blasted. Underneath each pocket is a vibrating grizzly feeder with the grizzly bars set at 4 in. (102 mm). These feeder-grizzlies are 84 in. (2. 13 m) wide and 32 .ft (9. 75 m) long, with grizzly sections 16 ft (4. 88 m) long. Approximately 75% of the ore passes through the grizzlies and grizzly oversize then drops onto two parallel 6 ft (1. 83 m) wide picking belts. The belts are 3 in. (76 mm) thick and rest on idlers supported by air- filled rubber shock absorbers. At the side of each picking belt is a hydraulic grapple which is used to remove timbers too large to be removed by hand. At times, four men are used to remove timber and tramp iron. Timbers are dropped onto a transverse belt conveyor which discharges into a truck stationed outside the primary crusher building,

Jan 1, 1978

By M. E. Wadsworth, W. H. Dresher, W. M. Fassell

HIGH temperature-high pressure techniques have long been used to great advantage in the organic chemical industry, the petroleum industry, and the paper industry. Only recently, however, have these methods been used to extract metals from their ores on a commercial scale. The Chemical Construction Corp., together with interested producers of nickel and cobalt, has done much to develop methods of producing nickel and cobalt powders by the use of high temperature-high pressure techniques.&apos; The Howe Sound Co. is currently operating a plant near Salt Lake City in which the new Chemical Construction Corp. technique is applied to cobalt-arsenic sulfide concentrates. In this method the concentrates are leached in a dilute solution of sulfuric acid under air pressure at a high temperature. Near Edmonton, in Alberta, Canada, the Sherritt Gordon Co. is also using the new method in conjunction with its ammoniacal leach process.&apos; National Lead Co., at Fredricktown, Mo., is operating a process similar to that of Howe Sound. In spite of the rapid growth of high temperature-high pressure hydrometallurgical processes, there has been very little work done on fundamental principles involved in this type of process. Conse- quently the literature lacks any significant data in this field. Oxidation of pyrite,3 sphalerite,4 galena," and recently molybdenite" in alkaline solutions has been reported to some extent. This investigation was initiated, therefore, to broaden the application of high temperature-high pressure hydrometallurgy and also to contribute some insight into the basic mechanisms of oxidation in an aqueous medium under these conditions. In addition to the immediate consequences of these processes much valuable information remains to be learned about fundamental reactions when they are carried out at temperatures and pressures slightly higher than those normally encountered. Leaching of molybdenite by alkaline solutions under oxygen pressure has recently been reported in the Russian literature by E. S. Usataya,6 ho found that leaching rate increased with an increase in the pH of the solution—the strongest effects being observed at a pH of 10—and also with an increase in temperature. Usataya also has reported the formation of a protective film on the surface of the molybdenite when the leaching solution was weakly alkaline, neutral, or acidic. He believes that this oxidation process explains the migration of Mo++ in the oxidation zones of molybdenite ore deposits and the subsequent impoverishment of the ores. Equipment: The high temperature-high pressure reaction unit used in this work was especially designed for kinetic studies of this nature, emphasis being placed on accuracy of measurement and precision of control. As details of the unit&apos;s construction and operation have been discussed elsewhere,&apos; only basic features and principles of operation will be discussed at this point.

Jan 1, 1957

By J. T. Shimmin

THE past twenty-five years has been a period of unparalleled improvement in all lines of mining and metallurgical equipment, but probably none has undergone greater refinement and development than the vacuum filter. Beginning a quarter of a century ago, there has been steady advancement in the art of vacu-um filtration until today its use has become virtually standard practice for dewatering throughout the entire metallurgical industry as well as in important parts of the chemical and industrial processing fields. The credit for the first successful application of vacuum filtration &apos;at the Consolidated Mercury mines in Utah in 1902 undoubtedly belongs to George Moore. There, it is said, his filter reduced the loss of gold in the slime and sand from $1.25 to 40c. per ton. Successful applications were subsequently made by Mr. Moore at the Liberty Bell mines, Telluride, Colo., the Lundberg, Dorr &Wilson mill at Terry, S. D., and elsewhere. E. L. Oliver is generally credited with, having done the most to work out the necessary continuity of op-eration exemplified in the well-known continuous vacu-um filter which bears his name. Others made improve-ments also, but in my opinion all the new ideas advanced in recent years are modifications of the fundamental principles originated by the two pioneers, Moore and Oliver. Considerable interest has been shown recently in the new type of continuous internal-type vacuum filter which is associated with my own name1. In response to frequent inquiries regarding it and the steps which led up to its invention, I am giving herewith a brief account of the early history of this development and some of the more important installations which have been made during the last year or so. A. M. Nicholas (now deceased), an Australian engi-neer, conceived the idea, several years ago, of manu-facturing a continuous vacuum filter with the filter medium applied upon the inside of the drum, as con-trasted with existing drum filters in which the medium is applied externally. Nicholas recognized that a ma-chine of this type would be simpler than one with a tank and would overcome some of the difficulties en-countered with tank filters on certain classes of pulps, and he appreciated the value of having filtration take place in the same direction as the gravitational settle-ment of the solid particles in the pulp, the pulp being retained within the drum instead of being held against the outside surface. The ease of starting up after a shut down was likewise realized by Nicholas. It was tried in the United States, but was not entirely success-ful primarily due to mechanical defects.

Jan 5, 1928

Copper ores from small mines in the area south of Tucson, where the Twin Buttes mine is located, were first operated in the 1880&apos;s. Today this area is one of the great copper mining centers of the world. In 1963, Anaconda Company acquired the property from Banner Mining Company and in recent years joined with AMAX in a 50/50 partnership to work the Twin Buttes property. From the beginning, Anaconda had relied on conveyors to remove overburden, oxide material for eventual leaching, and ore from the open pit. In the large st preproduction stripping program in mining history, Anaconda in a four-year period (1965 to 1969) removed 266 million tons of overburden of an average depth of 500 ft (152 m) lying over the ore body. The alluvium overburden was stripped by scrapers carrying 70 to 80 tons and push-loaded by two D9 tractors operating in tandem. The scrapers delivered their loads to bins which fed a 60 in. (1.52 m) conveyor having a capacity of 8500 tph. The 25% slope belt conveyor, which was 2100 ft (640 m) long, transported the material to the brow of the pit, where it was transferred to an 8300 ft (2530 m) surface belt to a 1000 ton loadout bin. At this point, the overburden was trucked to the sites of the tailings ponds, where it was used for dam construction. During 1976, the one billionth ton of overburden was removed from the Twin Buttes mine. The Twin Buttes quartz monzonite porphyry contains only scattered mineralization of which chalcopyrite is the most abundant mineral, associated with chalcocite, covellite, minor bornite, and several oxide minerals. Oxidation has progressed to variable depths of up to several hundred feet below the base of the alluvium. Because of the depth of the sulfide and oxide ores and hard rock waste, two in-pit crushers and conveyor systems were also installed for their transport out of the pit; and in 1976 a third crusher and conveyor system was installed lower in the pit for waste rock only. The two initial crushing and conveying systems were designed to handle any of three functions: crushing waste to be disposed of by a rail-mounted stacker -conveyor, crushing oxide ore for conveyor and then truck transport to stockpiles for future leaching, and primary crushing of sulfide ore for conveyor transport to the two 30 in. by 70 in. (0.76 by 1.78 m) secondary crushers and the two 7 ft (2. 13 m) standards and four 7 ft shortheads in the third and fourth stage crushing plant. A third standard crusher has been in- stalled recently. Mining is done by 15 yd (11.5 cum) shovels which load 100- and 150-ton

Jan 1, 1978

Discussion of the paper of LAWRENCE ADDICKS, presented at the Salt Lake meet-ing, August, 1914, and printed in Bulletin No. 91, July, 1914, pp. 1671 to 1674. JAMES H. PAYNE, Baltimore, Md. (communication to the Secretary*). -Mr. Addicks&apos;s account of the behavior of the nodulizer at Chrome on blast-furnace flue dust leads me to outline what has been done so far upon the application .of the same apparatus to fine sulphide concentrates. As Mr. Addicks states in his paper, the conditions prevailing in the test kiln at Yorktown in the demonstration I made for him upon Chrome material were strongly oxidizing; so much so that the sulphur was roasted down to less than 1 per cent. It is therefore to be expected that the first test made upon sulphide ore a short time after this, in which Ducktown green ore fines were used, did not show very promising results. Although the Ducktown ore contained only 13.52 per cent. of sulphur its gangue was quite fusible, and the intense heat generated by the oxidation of. the extremely fine sulphide particles slagged the mass, bringing about a condition approximating that of a Brückner roaster in distress. During the spring and summer of 1912, the writer kept working upon different schemes to cut clown this slagging. Two successful methods were finally worked &apos;out embodying opposite principles, one promoting oxidation of the finer particles and the other retarding. their oxidation. In the early part of the present year the writer succeeded in interesting the Braden Copper Co. in nodulizing, with the result that an extensive series of tests were made at Yorktown upon oil-floated concentrates covering quite a wide range in analysis. These tests led to trials in the Chrome rotary on the part of the U. S. Metals Refining Co. of Minerals Separation concentrates which they had in stock, and later to test runs to check up the Yorktown runs on Braden concentrates. The large-scale tests checked up with the Yorktown tests in every way (except fuel consumption, which was, however, correctly predicted), and led to the Braden company&apos;s decision to adopt the process at Braden. The fuel consumption upon oil-floated concentrates is not over 6 gal. per ton and in many cases is less. The action in the furnace is independent of the sulphur content, and there is actually less tendency to form nose rings in the Chrome rotary than upon flue dust. It is believed that the improvements entering into the design of the Braden kilns will cut down the nose-ring trouble to where it will no longer be a drawback to the process.

Jan 11, 1914

By A. G. H. Andersen, A. W. Kingsbury

SILICON bronzes containing iron are used to a considerable extent in industry, under the trade name of P.M.G. alloys. Various classes of wrought alloys fall in the composition range 1.5 to 3.5 per cent silicon, and 0.5 to 2.5 per cent iron. Castings may contain as much as 6 per cent silicon and 3 per cent iron at present. The constitutional diagrams of these alloys are thus of considerable practical importance. Copper-silicon-binary alloys have been thoroughly investigated by a number of workers, the latest reports being those of C. S. Smith1 and of A. G. H. Andersen.2 Although much work has been done on iron-copper alloys,3 recorded investigations on the copper-rich copper-iron binary constitutional diagram are not numerous. Tammann and Oelsen4 determined the solubility limits of iron in copper by means of magnetic measurements, and found that the solubility of iron in copper amounts to a fraction of one per cent at 400°C., increasing slowly to one per cent at 800°C. From 800°C., the solubility increases rapidly and attains 4 per cent at 1100°C. Hanson and Ford,5 using electrical resistivity measurements and the microscope, determined a somewhat different solubility limit. Of earlier X-ray works on the copper-iron system, very little has been published. The work by Bradley and Goldschmidt6 on copper-iron-aluminum alloys includes some information on the copper-iron binary. These investigators conclude that substitution of iron in the copper lattice decreases the spacings, which is contrary to the evidence of the present work. Hanson and West7 have investigated the ternary alloys of copper with iron and silicon. Although their diagrams are valuable contributions to the understanding of these alloys, they did not recognize the kappa phase,12, 7,8 and their diagrams do not everywhere conform to the phase rule. The present work contains a check by means of X-rays of the solid solubility limit of iron in copper, and an X-ray investigation of the ternary-phase diagram, which, owing to present conditions, it became necessary to curtail considerably. PREPARATION OF SAMPLES Castings weighing about 75 grams were melted under vacuum. The charges were made up of ingot iron, copper cathodes from Laurel Hill, and refined silicon supplied by the Electrometallurgical Co. Alundum crucibles were generally used. The exceptions were alloys 77 and 78, which were melted in graphite crucibles. The ingots were annealed at 800°C. for one week, then samples were cut for chemical analysis. Longitudinal sections of the binary alloy ingots were hammered into plates 3 in. thick. Wherever possible, these plates were homogenized at temperatures ensuring

Jan 1, 1942

By C. P. Kempter, H. L. Brown

Recently Chang determined heat content values of tantalum monocarbide and tungsten monocarbide from 325" to985°Kand 326" to 912"K, respectively, and, using other published data, made certain solid-state calculations for these two compounds. Chang expressed his Cp values in units of cal per "g-atom alloy" OK. We consider this poor usage since l) tantalum monocarbide and tungsten monocarbide are not alloys, and 2) an atom of an alloy does not exist. However, one may convert his values to units of cal per mole OK by merely doubling them. For TaC he calculated a Nernst-Lindemann parameter of 9.07 x lo-&apos; "g-atom per cal". We have refined the calculation of this parameter and have arrived at a much lower value. Rather than approximating the true coefficient of volume expansion, a,, as three times the true coefficient of linear expansion, a, we calculated u, from the TaCo.99, lattice expansion data of Fries and Wahman.38 At 25°C a, = 14.3 x 10s per OC. The molar volume was calculated to be 13.32 cu cm per mole at 25°C. From the lattice compression data of Champion and Drickamer~&apos; for "TaC,.,," we determined the room-temperature compressibility to be 3.5 x 10-l3 sq cm per dyne by plotting l/Vo (aV/ap) vs applied pressure. (Since the lowest pressure at which Champion and Drickamer reported a relative volume was 20 kbars, one may calculate an approximate compressibility only. Chang calculated a compressibility of 3.175 x 10-l3 sq cm per dyne from the same data.) We also used twice Chang&apos;s value of Cp = 4.38 cal per "g-atom alloy" OK for the molar heat capacity at 25°C. From these four values we calculated a Nernst-Lindemann parameter of 5.8 x 10-l4 mole per erg or 2.4 x lo- &apos; mole per cal, and a Cdil value of 2.3 x 10&apos; ergs per mole OK. In the case of tungsten monocarbide no isothermal compressibility has been reported; however, one may calculate an approximate Nernst-Lindemann parameter by using the adiabatic compressibility reported by Brown, Armstrong, and Kempter." Using the WC lattice expansion data of Mauer and Bolz~l we calculated a, to be 12., x l0-&apos; per OC. The molar volume was calculated to be 12.5 cu cm per mole. We used twice Chang&apos;s value of Cp = 4.18 cal per "g-atom alloy" OK for the molar heat capacity at 25°C. From these four values we calculated an approximate Nernst-Lindemann parameter of 2 x lo-&apos; mole per cal, which is much

Jan 1, 1969

By B. H. Caudle, A. B. Dyes, R. A. Erickson

The study of the influence of fluid mobilities on the sweepout pattern resulting from the injection of gas or water has been extended to cover the production period which follows breakthrough of the injected material. Mobility ratios over the range common in field operations (0.1 to 17) were studied for several pattern floods (five-spot, staggered, and direct line drive). The experimental data required for these studies were obtained by the use of the x-ray shadowgraph technique using miscible oil phases of different viscosities in porous plate mode1s of a reservoir element. From the shadowgraph pictures obtained before and after breakthrough of the injected fluid, flowing ratios at the producing well and cumulative volumes injected were calculated. The method for applying such data in predicting field behavior is illustrated for a water flood of a five-spot. For this case a range of mobility ratios of 0.5 to 5.0 results in: (I) nearly complete (95 to 100 per cent) sweepout pattern efficiencies at abandonment conditions, (2) production after breakthrough being responsible for as much as one-third of the total recovery at the lower mobility ratios, and (3) a twofold variation in the operating life of the reservoir. In determining the merits of a secondary recovery operation, the fraction of the reservoir which is ultimately to be swept by the injected fluid is of major importance. Because of its significance a large amount of research has been applied to the study of areal sweepout efficiency. Muskat, Wyckoff and Botset1,2 have presented the classic analysis of the pattern efficiency at breakthrough for several injection patterns in which the mobility of the fluids ahead of the injected phase front was equal to the mobility of the fluids behind the front. Aronofsky3 as extended this analysis of recovery at breakthrough through numerical computations and by the stepwise use of the potentiometric model to include the effect of different fluid mobilities ahead of and behind the invasion front on the line drive performance. Fay and Prats&apos; have also employed numerical computations to show the influence of one other condition of fluid mobilities on the five spot flood. In a previous paper&apos; a new method of attack on this problem which employed x-ray and porous plates was described. At that time experimental results were also presented which described the effect of fluid mobilities on the reservoir area contacted by the injected fluid at breakthrough for the five spot and direct line drive patterns. Most of the literature pertaining to sweepout pattern efficiency has emphasized the pattern which is obtained at breakthrough of the injected phase. The increase in the sweepout pattern which occurs after breakthrough

Jan 1, 1955

By Walter E. Duncan

THE ores treated at the concentrating plant largely come from the blanket replacement deposits of the northeastern part of Hardin County, Illinois, and consist of complex mixtures of galena, sphalerite and fluorspar in a gangue of quartz, limestone and sandstone. The several minerals are so intimately intergrown that they are not liberated at the sizes usually treated by gravity concentration in the usual fluorspar plant. For this reason, an all-flotation method of concentration was developed in cooperation with the station of the U. S. Bureau of Mines at Rolla, Mo.1 This flotation separation made possible the commercial production of four concentrates: (I) lead concentrates, (2) zinc concentrates, (3) acid-grade fluorspar concentrates and (4) metallurgical-grade fluorspar concentrates.2,3 Although fluorspar had been floated in Canada in 1921 4 and a plant at Rosiclare had been floating fluorspar since March 1929, fluorspar flotation was given considerable impetus by the Mahoning developments, since some features of the separation were quickly adopted by other fluorspar companies. The Bureau of Mines5 reports that in the 10 years preceding 1938 only 50,552 tons of fluorspar flotation concentrates were produced in the United States. The rise from this almost nominal figure to 153,750 tons (largely acid grade) in one year (1944)6 is indicative of the activity in fluorspar flotation occasioned by the war. The industry thus proved to be in a position to expand and keep up reasonably well with the requirements imposed upon it by such developments as occurred in high octane aviation gasoline alkylation, in the use of Freon as a refrigerant and as carrier in the Aerosol "bug bomb," and in the atomic bomb project as well as in the expansion of other activities requiring hydrofluoric acid; such as, for example, the aluminum industry. Despite the fact that the greater portion of acid-grade fluorspar produced in the last few years went directly into war uses, the future of the acid-grade field still appears bright. As a result of the activity in fluorine chemistry during the war, it seems likely that in the relatively near future almost every phase of human endeavor will find some fluorine compound involved. The other three concentrates produced deserve some attention in these introductory remarks. The lead concentrates first produced contained both zinc and fluorspar but these did not interfere with their marketing. As will be pointed out later, the quality has steadily improved until today it is a relatively high-grade product. Although some zinc smelters accept concentrates containing limited quantities of fluorspar, it is a very objectionable impurity. Since flotation methods fail in the complete removal of the fluorspar, chemical methods were developed for

Jan 1, 1946

By E. W. McAllister

THE decision of the California Oil Umpire&apos;s1 office to accept well potentials established from subsurface-pressure data has brought to the attention of many operators for the first time the application of such data to production problems. This paper reviews the development and application of subsurface-pressure data in Kettleman Hills, where these data have been used in production work since 1932. The Kettleman Hills field was discovered in October 1928. The first subsurface-pressure surveys were made early in 1932 and by the middle of 1933 the two major operators in the field were making routine pressure runs in their respective wells. Shortly thereafter, these companies agreed to exchange pressure data, and later a third company entered into the agreement. This made it possible to analyze the effects upon the reservoir pressure of past production practices, which are discussed in the latter part of the report. It is interesting that many of the applications of subsurface-pressure data discussed herein were anticipated by C. L. Clarks and C. B. Kimberlin‡ in a reports dated March 1933. PRODUCTIVITY INDEX In the majority of individual well problems involving the use of subsurface pressure data, it is necessary to use a factor known as the "productivity index," which indicates the ability of a well to produce and represents the number of barrels of oil or oil and water per day that will enter the well bore for each pound per square inch pressure drop made opposite the producing interval of the well. The productivity index may be determined from pressure and production data for one or more rates of flow and a static or shut-in pressure survey or from production and pressure data for two or more rates of flow. Productivity indices may be based upon either net oil or total fluid production. All indices referred to in this report are based upon net oil. The value of the productivity index for Kettleman Hills wells decreases generally with time and it is not uncommon for the index to vary in value at any specific time for various rates of flow. The decline of the productivity index may be attributed partly to the loss of reservoir pressure. As. the reservoir pressure decreases, gas originally in solution is released from the oil, congesting the flow of fluids through the formation and increasing the viscosity and surface tension of the oil; thus the flow resistance of the fluids through the formation is increased, resulting in a decrease of the productivity index. For convenient reference, the productivity indices for each well are plotted against time and a curve is drawn through the points showing the trend of the average index value. Plotted on Fig. I are index values for five

Jan 1, 1941

By R. W. Jordan, J. Winter, A. J. Goldman

Kirkendall diffusion of copper and nickel atoms across the bond interfaces, as well as oxygen diffusion in the core to the interfaces, are the mechanisms which control the variation in bond strength with annealing time at 1600°to 1800°F in the cupro-nickel clad copper Mint composite. The unequal mass transfer resulting from the high diffusion rate of copper in cupro-nickel compared with that of nickel in copper caused void formation in the core. Small voids also formed in the cladding after short annealing times. These may have been nucleated by nickel oxide particles which are formed by the dissociation of the cuprous oxide in the core and the movement of oxygen to the cupro-nickel interfaces. The bond strength drops to very low values for annealing time up to 3 hr when the void density is greatest. The increase in bond strength after 16 to 72 hr at 1600°F is a result of the decrease in the void density in the cupro-nickel cladding and the growth of grains across the interface. The void density decrease is attributable to grain boundary diffusion of vacancies to the surface. The abnormally large grains in the copper preclude a similar process from taking place in the core. THE solution to many engineering materials problems that cannot be satisfied by using a single metal or alloy is often attained by combining two or more metals in such a way that each component performs a different function, while the composite meets the overall requirements. The Coinage Act of 1965 authorized a changeover from the traditional 90 Ag-10 Cu dimes and quarters to a bimetallic laminate of 75 Cu-25 Ni clad on a copper core. Economic considerations as well as physical, chemical, and mechanical properties provided a basis for the selection of this composite as a substitute for the silver-bearing alloy. An extensive survey of alloys suitable for use as coinage was prepared by Battelle Memorial Institute for the U.S. Mint.1 It was found that the cupro-nickel/cop-per/cupro-nickel composite satisfied the required color, density, mechanical properties, corrosion resistance, electrical properties, and magnetic properties to make it completely acceptable to the general public and the vending machine industry. This investigation is part of a program to determine the feasibility of reclaiming composite scrap by thermally degrading the bond strength of the cupro-nickel clad copper laminate. The composite exhibits bond degradation and subsequent bond improvement during annealing in the temperature range 1600" to 1800°F. A study of the operative diffusion processes and the resultant structural changes was made in order to explain the observed variations in bond strength with annealing time. EXPERIMENTAL PROCEDURES AND RESULTS Bond Degradation. A measurement of the forces required to peel the cupro-nickel cladding from the

Jan 1, 1969

By C. W. Haworth, M. G. Hall

THE diffusion of copper in silver has been studied by Sawatzky and Jaumot1 and Cahoon and Youdelis.2 Sawatzky and Jaumot measured tracer diffusion in single crystal specimens but the 12.9 h half-life of CU64 limited the time of the diffusion anneal to about 50 hr, and in consequence the minimum temperature of the investigation was 716°C. Cahoon and Youdelis employed polycrystalline diffusion couples, and measured the interdiffusion coefficient over the range 0 to 2 at. pct Cu. These experiments were performed at temperatures down to 630°C and a discontinuity in the Arrhenius plot was observed at -700°C. The high D-values at lower temperatures were attributed to the effect of grain boundary diffusion. The present experiments were undertaken to investigate this anomaly and study the low temperature region of the diffusion of copper in silver. The method employed single crystal samples and a plated source of copper. Silver single crystals 2½ by 4 in. sq were grown from zone-refined silver bar (>99.99 pct Ag) by the Bridgeman technique, and $ in. thick slices were cut by spark machining. The specimens were electro-polished, and a 2 µm layer of copper was electroplated on to the planar surface at one end of each specimen using a potentiostatic voltage source. The following solution was used for plating: CuSO4 . 5H2O 125 gm H2SO4 (S.G. 1.8) 53 ml H2O 1000 ml The diffusion couples were sealed in capsules, in an argon atmosphere, and were annealed at either 701, 597, or 498°C. The temperature measurements were made to an accuracy of *l°C After the anneal, the penetration of copper into the silver was determined using an electron probe micro-analyzer. 100 sec scaler measurements of the intensity of Cu Ka! X-radiation were performed at 2 to 5 µm intervals along a line normal to the source, with a beam potential of 25 kv. Allowances were made for the background radiation, and the dead-time of the counting system in the normal way, and the intensity ratios were corrected to weight concentration by the method of Duncumb & de-Casa.3 For the analysis of the diffusion profiles, it was first observed that the copper layer did not completely dissolve in the host crystal for the specimens annealed at 498 and 591 whereas for the specimen annealed at 701°C, a single phase solid solution was formed. For the 498" and 597°C specimens, the concentration of copper in the silver at the interface was therefore constant throughout the anneal and the analysis for a

Jan 1, 1970

One of the best known of the early mining districts of Nevada was that at Eureka, where gold was first produced in 1864. While the most productive period for the district was during the early 1900s, production was maintained until World War II. During this time it is said that the Eureka district produced as much as $200 million from over 3,200 km (2,000 miles) of tunnels and workings. The Windfall Mine was discovered in 1902; and a 100 ton cyanide plant was operated there during the periods 1907 to 1912 and 1932 to 1940. Since 1974, however, Windfall Venture has been operating a unique heap leaching facility on the site of the old Windfall Mine. Gold at the Windfall occurs primarily in the Cambrian Hamburg dolomite which has been "sanded" and in which the original cementing material has been destroyed. The gold is micron sized and is accompanied by anomalous amounts of arsenic and mercury; however, the ore is low in base metals which distinguishes it from the typical Eureka district ores. Mining the ore causes it to break down into very fine sizes, approximately 50% minus 150 microns (100 mesh). Conventional heap leaching by spraying proved to be impossible; and extensive test work on alternate methods, aided by the United States Bureau of Mines, was conducted using 9 m (30 ft) high columns and then large tonnage heaps. The leaching procedure adopted was to build 10.6 m (35 ft) high heaps on a 50 mm (2 in.) asphalt pad covered with a layer of gravel for drainage assisted by perforated pipe drains and gravel-packed sumps. A berm is constructed around the top of each pad and cyanide solution is allowed to “pond” on top of the heap. The resulting percolation is at the rate of 24 liters per sq m per hr (0.6 gal per sq ft per hr), a figure about 2.4 times the average used in other heap leaching operations. The dolomite furnishes its own protective alkalinity and no lime or caustic is used in the solution which contains 0.4 kg per mt (0.8 lb per st) of NaCN. Pregnant solution from the heaps averages about 0.7 g Au per mt (0.02 oz per st) with a cyanide content of about 0. 2 kg per mt (0.4 lb per st) and a pH of plus 9. Addition of NaCN brings the pH of the leach solution up to 11 to 12. Carbon adsorption of the gold was originally done in six enclosed columns mounted in a trailer but five conventional columns have been substituted as the size of the operations has increased. Carbon stripping is with hot caustic, cyanide and 10% ethanol solution. The

Jan 1, 1981

By H. B. Aaron

DIFFRACTION micrography provides a useful tool for studying complex strain fields. Newkirk1 observed an X-ray diffraction effect due to strain interactions at the intersection of slip lines in LiF and attributed this effect to a reduction of primary extinction due to strain-gradients. It is the object of the current study to correlate the observations and rationalize the appearance of the strain fields around intersecting slip planes in x-ray extinction contrast studies by a semiempirical method. Fig. 1 shows typical strain patterns observed by Newkirk on (200) reflections in LiF using chromium radiation and a variation of the Berg-Barrett (B-B) technique. In other reflections (e.g., {220}) the strain patterns are absent. If one defines a vector 3 along the striations from the more to the less dense portion, and a vector v in the direction of the X-ray beam parallel to the reflecting plane, such that both 3 and v lie in the reflecting plane of the crystal, Fig. 1, then the positions of the striations on (200) reflections appear to be confined such that where is the angle between v and 3, Fig. 1. It will be seen later that 3 is the projection of Sb on the reflecting planes. In some instances the strain effects are observed only above the slip plane, Fig. l(C), in others only below, Fig. l(D). Where the strains appear to be both above and below the slip line, Fig. l(E), one sees that there are actually two slip lines present, Fig. 1(F) and (G). Furthermore, the strain patterns are observed in the vicinity of the intersection of emergent edge dislocations (diagonal lines on the micrograph) with emergent screw dislocations (horizontal lines on the micrograph), e.g., Fig. l(HJ) and (MN), but not in the vicinity of the intersection of two emergent edge dislocation arrays, Fig. l(HJ) and (KL). However, in the latter case the contrast of the dislocation array itself is reduced along (KL) and increased along (HJ). Line (AB) in Fig. 1 is the edge of the crystal, and the material to the left of (AB) is the side of the sample. Any empirical correlation will have to include most, if not all of the aforementioned observations. Lang2 and Newkirk1 determined that the condition for contrast is given by g-b#0 [2] where b is the Burgers&apos; vector and g the diffraction vector (normal to the reflecting planes). If one assumes that the direction of the observed strain fields is given by the vector sum of the slip vectors (Sb1) of the intersecting slip planes, then Eq. [2] may be generalized to g-S#0 [3]

Jan 1, 1963

By Reuben B. Ellestad, Fremont F. Clarke

Lithium chemical plant extraction methods are discussed with reference to 1) base exchange with alkali sulphates; 2) processing based on roasting with lime; 3) miscellaneous methods; and 4) application of the Lithium Corp. process to extraction of lithium from run-of-mine, low grade spodumene ore, or concentrates. IN the early days of the lithium industry most of the production was from lepidolite, zinnwaldite, and amblygonite. Nearly all the early extraction processes described in the literature involve heating the finely ground mineral with sulphuric or hydrochloric acid. On subsequent water leaching most of the bases present in the mineral (especially aluminum) are dissolved as sulphates. As a result, the leach solution required extensive chemical purification before the lithium could be precipitated as carbonate. Following the remarkable growth of the lithium industry to its present size, zinnwaldite and amblygonite ores must be considered of minor importance only. Attention is now focused on spodumene, abundant enough in North America to be a major source of supply, and there are important supplies of lepidolite and petalite in Africa. The extraction processes described below all apply to spodumene, although several will also operate on other lithium minerals, such as petalite. Base Exchange with Alkali Sulphates: A distinct advance was made with the disclosures of Wadman and von Girsewalt. In these methods the finely ground silicate ore (spodumene or lepidolite) is intimately mixed with an excess of alkali sulphate (usually K2SO4) in at least a 1 to 1 proportion, and the mixture was heated to a relatively high temperature. Base exchange results, with the formation of lithium sulphate. A water leach dissolves the lithium sulphate, together with the excess potassium sulphate. Successful operation of this type of process requires very thorough grinding and mixing, as well as careful temperature control. The use of K2SO4 is objectionable from cost considerations since purification of lithium carbonate requires the use of potassium carbonate, if the K2SO4 is to be recovered and recycled. The lower solubility of K2SO4 as compared with Na2S04 is also objectionable, since it limits the concentration of the Li2SO4 solution to be precipitated by K2SO4. Early laboratory-scale investigation of this process by Lithium Corp. was not encouraging. Other related base exchange processes are those of Lindblad, Wallden, and Sivander3 and Sivander, Gard, Villestad, and Wallen4. The former covers the reaction of lithium silicate minerals with a sodium sulphate solution, at 100" to 300°C (under pressure), while the latter involves the extraction of silicate minerals with molten sodium sulphate. Both these processes would seem to be difficult and expensive to operate.

Jan 1, 1956

By N. P. Peterson

The Castle Dome copper deposit is of the porphyry type and occurs in a body of quartz monzonite intruded into the pre-Cambrian formations and possibly into the lower part of Paleozoic limestones. The hypogene sulphide minerals include pyrite, chalcopyrite, and a little molybdenite, sphalerite, and galena. Pyrite and chalcopyrite are disseminated or occur in small closely spaced veins. The hypogene mineralization shows distinct zoning in the distribution of pyrite and chalcopyrite, so that the mineralized area can be subdivided into pyritc and chalcopyrite zones. Supergene enrichment, though not extensive, played an important part in the formation of the ore body, the chalcopyritc zone being the pro-tore. Several thin, gently dipping, diabase sills occur in the mineralized area, and the copper content of the rock in and near these sills is substantially higher than average. The formation of the ore body was caused by the combined effects of three forces: I. Zoning of the hypogene mineralization, 2. Richer copper metallization associated with the diabase sills, and 3. Supergene enrichment. Introduction The Castle Dome copper deposit is about 5 miles west of Miami, Arizona, and about 3 miles west of the Miami-Inspiration ore body. The Castle Dome open-pit mine is on the south flank of Porphyry Mountain, a prominent landmark of the region. In 1941 the deposit was being explored by the Miami Copper Co., but the exploration program was cut short when the decision was made to develop the property as a war project. The Castle Dome Copper Co., Inc., was organized as a wholly owned subsidiary of the Miami Copper Co., and arrangements were made with Defense Plant Corp. for developing and equipping the mine and for construction of a concentrating plant. Construction and development work was begun in January 1942, and milling operations started June so, 1943. Since that date, the mine has been in continuous operation by the Castle Dome Copper Co, under a lease agreement with Defense Plant Corp. producing copper at an average rate of 4,000,000 lb per month. To January I, 1946 the total production amounted to 121,088,000 lb of copper,4,271 oz of gold, and 211,755 oz of silver. A detailed study of the copper deposit and the surrounding area was undertaken by the U. S. Geological Survey in June 1943. Other members of the Survey participating in the work were C. M. Gilbert, G. L. Quick, and J. P. Albers. The results of this work will be published as a comprehensive report by the Geological Survey. General Geology The Globe quadrangle, which includes the Castle Dome area, was studied by F. L. Ransome1 of the Geological Survey in 1901 and 1902. His maps and descriptions were

Jan 1, 1949

By W. H. Herrnstein, F. H. Beck, M. G. Fontana

Sulfur solubility in a iron was measured in the 750 to 890°C temperature range and was found to be insensitive to ferrite purity within a limited composition range. The permeability of sulfur in ferrite was determined from the internal sulfidation of Fe-Mn alloys. From the solubility and permeability results, the diffusivity of sulfur in ferrite could be calculated. These results are summarized as follows: Dissolved sulfur decreased the lattice parameter of ferrite. The effect of dissolved sulfur on the tensile properties of ferrite at low temperatures was also determined. SINCE sulfur is an important residual impurity in commercial irons and steels, effects associated with the presence of sulfur as a solute in iron are of considerable practical interest. Unfortunately, certain basic information relative to the Fe-S system appears to be lacking in the literature, e.g., the effect of dissolved sulfur on the lattice parameter and mechanical properties of ferrite. In addition, discrepancies exist in previously reported results for the solubility of sulfur in ferrite. These discrepancies raise questions regarding the sensitivity of sulfur solubility to the purity and microstructure of iron. This study was undertaken primarily in an effort to obtain reliable data on the solubility of sulfur in ferrite. Additional objectives included determination of the permeability and diffusivity of sulfur in ferrite by the internal sulfidation method and determination of the effect of dissolved sulfur on the lattice parameter and tensile properties of ferrite. EXPERIMENTAL METHOD Three grades of iron, encompassing a range of impurity levels, were employed in the study. The grades were designated VME (vacuum-melted electrolytic), SPZR (single-pass zone-refined), and MPZR (multipass zone-refined). The composition of each grade is given in Table I. Poly crystalline and single-crystal strip specimens were employed in the solubility experiments. Uniform grain size strip (ASTM 3 to 5) of 0.005 to 0.020 in. thickness was obtained by alternate cold reductions of 60 thickness 60 to 80 pct and recrystallization at 650°C for i hr in pure, dry argon. Single-crystal specimens were produced in 0.010-in.-thick SPZR iron strip by the strain-gradient anneal technique. Recrystallized strip 0.25 in. wide was strained until a Luders band or bands had propagated over the length of the strip; the strip was then annealed by drawing it through a 150°C per in. gradient (890°C peak temperature) at 1 in. per hr in an atmosphere of pure, dry argon. Laue photographs of the crystals produced exhibited no discernible asterism. Specimens were equilibrated with H-H, atmospheres at temperatures between 750" and 890" by one of two methods, each of which makes use of the reaction: The encapsulation sulfurization method was employed in the majority of the solubility experiments. A strip

Jan 1, 1969