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By Virgil Miller, F. B. Petermann, S. F. Ravitz

The Cle Elum nickeliferous iron deposit is in Kittitas County, Washington, in a rugged, mountainous region about 23 miles north of the town of Cle Elum. The Biewett Pass deposit, which is similar in character and occurrence, is in Chelan County, about 18 miles east of the Cle Elum deposit. A more or less continuous ore zone may extend between the two deposits, since iron ore is reported to have been discovered at Iron Mountain, which lies directly between them. The ore probably was formed by meta-morphism of laterite resulting from Ieach-ing of serpentine derived from peridotite. It consists largely of magnetite grains distributed through a matrix of limonitic material containing a small amount of hematite and chromite. Many of the chromite grains have shells and inclusions of magnetite. The magnetite grains range in size from 1200-mesh to 1/4-in., the average size being 48 to 65-mesh; the chromite grains range from 20 to 150-mesh. The gangue minerals are olivine and serpentine with a small, amount of chlorite, and appear both as very thin bands in the ore and 's minute fragments in the limonitic matrix. The nickel is present as minUte inclusions of a nickel silicate mineral within the magnetite and gangue par- ticles, varying in size from 1500 to about 280-mesh. Detailed analyses of 3000-lb. lots of Cle Elum and Blewett Pass ore are given in Table I. TABLE i.—Analyses of Ore Samples Per Cent Ore from Constituent cle Elum' Blewett Pass Cr....................... 2.23 2-12 Ni...................... 1.43 0.9s Fe...................... 52.9 41.4 SiOz.................... 5.6 13.4 CaO.................... 0.3 0.8 MgO................... 4-1 7-5 AI2O3................... 7.0 2.g P...................... 0.022 0.034 S....................... O.O1 o.oa Mn..................... I.I 0.4 Ba..................... nil co..................... o.oa 0.02 CU..................... 0.02 0.03 As...................... 0.03 Au..................... nil Ag°..................... O.58 Pt...................... ml Fe++.................... 8.9 5-9 Fe+++.................. 44-0 35-5 « Ounces per ton, 1 sr,cte,so,","&&$ally absent: B, W, M~, T~, V, ~i, Bi, Pb. Sb. Sn. Be, In, Hg. Zn, Zr. Cd. Preliminary Tests Attempts to beneficiate the ore by ore-dressing methods, including various 'cornbinations of sink and float, tabling, jigging, flotation, and magnetic separation, were consistently unsuccessful. A wide variety of roasting and leaching tests were made on both raw and reduced ore, but in most of the tests little or no nickel was extracted, and in none did the extraction exceed 50 Per cent. No nickel could be volatilized as the carbonyl. by treating reduced ore with carbon monoxide. Fairly good sepa-

Jan 1, 1944

By M. A. Rigdon, R. E. Grace

The dissociation of cupric oxide to cuprous oxide and oxygen was studied with a microbalmce technique at 700" to 750°C. In this temperature range the dissociation pressure of the reaction 2CuO= Cu2O + 1/2 O2 was: The theory of near-equilibriwm reaction rates predicts a linear relationship between reaction rate and the Gibbs free-energy change for small values of AF/RT. This correlation is found to be valid for AF/RT = -0.1 or approximately 200 cal per mole away from equilibrium. THE theory of reaction rates near equilibrium as developed from classical thermodynamics suggests a small range of linear behavior between the overall reaction rate and the free-energy change accompanying the process. The extent of this linear correlation is generally unknown, even to the order of magnitude. The idea of linearity is also found in the formal mathematics of coupled transport phenomena, e.g., chemical diffusion, thermal diffusion, thermoelectricity, and so forth. The linear array of forces and fluxes which leads to various main and cross coefficients is predicated upon linearity between reaction rates and the bulk thermodynamics which govern the process. The general case in reaction kinetics is well-known: the magnitude of the free-energy change does not cor- relate with observed reaction kinetics, irrespective of whether the process occurs by diffusion or chemical reaction or both. The real need is to investigate specific reaction paths near equilibrium in order to learn of the limitations in attempting to correlate reaction rates and thermodynamics. It is the main purpose of this paper to examine the dissociation kinetics of cupric oxide under conditions of known driving force where the dissociation process is likely to occur by chemical reaction. A secondary objective is to obtain equilibrium data for the CuO-Cu20-O2 system. In a prior paper Edmiston and race' investigated reduction kinetics of wustite in H2-H2O gas mixtures. The present study was carried out in oxygen according to the following reaction:

Jan 1, 1969

By Morris Cohen, Robert C. Ruhl

The phases in Fe-C alloys over a wide composition range have been studied after splal quenching from the liquid state. Binary alloys containing 0 to 5.1 wt pel C as /cell as a large number of ternary Fe-C-Si alloys with 2.5 to 5.0 wt pct C and 0.3 to 5.1 wt pct Si were attlong those sludied. Olher Fe-C-X alloys, zcilh X being Co, Cr, ,Wn, Ni, and Ru , were also inrestigated after splat quenching. At high carbon contents, a new hcp phase (designaled 6 phase, but different from e carbide) is retained upon splat quenching. The .fraction of this phase varies up lo 97 pcl for a Fe-4.K-1.9% alloy. The composition of the E phase ranges from about 3.8 10 4.8 wt pet C, and the corresponding laltice parameters increase linearly ulith carbon content, while the c/a ratio remains essentially conslc~nl. The E phrtse appears to he a solulion of carbon in E iron, the latler being nornially found only at high pressures. It is deduced that the unit cell of the E phase corresponds to the formula Fe12C3, and llzal il is relaled to tlie ordered slruclures 0.f 6 iron carbide and c iron nilride. The E phase is compared and contrasted to the olher known carbides and nitrides of iron and nickel. An exlrapolaTion of the atomic volume 1,s carhon conlent of /he E pllase lzcts giz.en a neu7 estitnale jor /he alomic volume of E iron, 11.30 cu A, a1 atmospheric pressure and temperature. Other alomic volume relalionships lead to /he co~zpositioti Fe2.iC tor E iron carbide, /he unit cell fortr~ula being -Fe2rClo The E phase undergoes a lulo-slage decomposition upon healing, .forrning firsl rnarlensile plus E carhide, a/ler 1 hr at 140" lo 200°C, slid then ferr ite Plus cementile, after 1 lir a1 330" to 460°C. A1 carbo,l contents between 1.5 and 3.0 LC/ pcl, (he predo.wirzar/t plzase alley quenching is fcc austenite. The retained carbon content of this phase increases with itlcreasing silicon in certain concentration ranges, reaching a maximum of 2.37 wt pct C itz a Fe-2.6C-4.OSi alloy. This is the highest carbon conten1 reLaitled in austenite to date. These high-carbon aus-tenites can be partially tm?zsforttled lo tnartensile hy severe deformation in the temperature range of — 190 to -50°C. TECHNIQUES for splat quenching from the liquid state have been utilized in numerous recent investigations to produce metastable phases in a variety of alloy systems. Among the several ways of splat quenching, the shock-tube method appears to yield the highest cooling rates1-3, 7, 8 and was adopted here. Estimated cooling rates attained in the present experiments ranged from 10' to 10 80Cper sec.' As a part of a research program on interstitial al- loy phases, the Fe-C system was selected for splat-quenching studies. It was hoped that splat quenching would allow high metastable supersaturations of carbon to be retained in solution. Also of considerable interest were the conditions governing the occurrence of the various intermediate phases upon solidification. The alloys investigated included both binary Fe-C compositions as well as six ternary Fe-C-X alloy systems. The known phases in the Fe-C system are summarized in Table I.* Only the ferrite and graphite are ent investigation.14 and is described in detail herein Table II summarizes corresponding data on Fe-N phases, which are also of interest here because of their similarity to the Fe-C phases. EXPERIMENTAL PROCEDURES Alloys were prepared by melting the elements. 99.9 pct purity, in an inert gas nonconsumable electrode arc furnace. The buttons. weighing about 5 g. were remelted twice. were then fragmented. and their interior surfaces were examined for uniformity: if any doubt existed. they were remelted again. Chemical analyses were performed on all the alloys. the accuracy being about k0.05 wt pct. At high temperatures, carbon-containing alloys react with alumina crucibles as follows: If the atmosphere in the splat-quenching furnace does not contain sufficient carbon monoxide the alloy can be depleted of carbon and contaminated with aluminum. Calculations and experimental observations showed that 50 to 100 torr CO partial pressure effectively blocked the above reaction in all the alloys investigated. The splat-quenching equipment in Figs. 1 and 2 provides for evacuation and back-filling with CO-Ar mixtures. The furnace is capable of operation up to 1650°C. and the gettering action of the graphite heating element reduces the oxygen partial pressure in the furnace atmosphere to below 10-5 torr, thus preventing oxidation of the specimens.

Jan 1, 1970

By S. E. Szasz, E. J. Moore, B. F. Whitney

An accurate value of formation water resistivity R, is essential in calculating formation porosity and fluid saturation from electrical well logs. In the cases where R, has not been measured directly, it must be obtained from other data, e.g., the SP curve. This paper deals with another approach: how to calculate R, from the chemical analysis of the formation water. INTRODUCTION It is known that the resistivity of aqueous solutions of pure salts depends on their concentration and on the temperature; the concentrations are given in MPL (mg of solute per liter of solution), or sometimes in ppm (mg of solute per kg of solution): MPL = ppm X specific gravity. Values for different pure salts are available in the literature, but not for solutions of mixtures which are of practical interest. The major component of the dissolved material in almost all formation waters being sodium chloride, it is customary to express the resistivity of formation waters in terms of equivalent sodium chloride concentration, i.e., the concentration of a solution of pure NaCl which has the same resistivity at a given temperature as that of the formation water under consideration. Thus, the problem of calculating R, from the chemical anaylsis can also be stated as how to convert the other constituents of the solute into equivalent NaCl concentration. Salts dissolved in water are at least partly dissociated into ions, and do not conserve their identity. If known amounts of several salts are dissolved in water, the solution does not necessarily contain the same salts in the original proportion, but perhaps some other combination of the ions, along with free ions in solution. This is why the chemical analysis of formation waters is often given in terms of ions, as if all dissolved salts were completely dissociated. Our problem then boils down to how to convert the concentrations of the various ions to equivalent concentrations of Na' and C1-. Dunlap and Hawthorne' have proposed to convert the concentration of all other ions to equivalent Na' and C1-concentrations by means of constant multipliers; e.g., 0.95 for Ca"; 2.0 for Mg"; 0.27 for HCO 3-; 0.5 for SO, -, etc. Their factors were based on measurements made at 68F on 26 formation water samples from the Texas Gulf Coast, ranging in concentration from 1,500 to 75,000 ppm. The Dunlap method is widely used in electric log interpretation, and is often extrapolated beyond its original concentration range. A comparison of R, values calculated by this method and values actually measured on formation water samples has shown large discrepancies, especially at higher concentrations. Therefore, two new methods were developed at Sinclair Oil Corp.'s Tulsa Research Center to calculate equivalent sodium chloride concentration from the chemical analysis of formation water samples. FUNDAMENTAL CONSIDERATIONS The resistivity of a solution, or its reciprocal the conductivity, at a given temperature is determined by the charge, concentration and mobility of the ions actually present. Monovalent ions such as Na' or C1- always carry the same charge. Compounds of polyvalent ions, however. may show incomplete dissociation, e.g., NaSO; + Na' instead of SO,-- + 2Na'. This happens especially in more concentrated solutions. Only very dilute solutions are completely dissociated, as assumed in the chemical analysis report. At higher concentration, the degree of dissociation depends not only on the nature and concentration of the particular salt under consideration but also on the nature and concentrations of the other solutes. Mobility of the ions depends on the viscosity of the solution. It also depends on the degree of hydration of the ions, which in turn is a function of the nature and the charge of the ions and also of the amount of free water available per ion, i.e., the total ionic concentration. The net effect is that the conductivity increases slower than proportional to the concentration, even if a solution contains only one salt such as NaC1, and is different for different salts (Fig. 1). Conductivity can even decline with a further increase in concentration, e.g., if additional salt is little dissociated but ties up some of the free water and/or causes an increase in viscosity. In solutions containing more than one salt, the contribution of one salt to the total conductivity depends not only on the fractional concentration of this same salt, but also on the concentration of all other solutes. A perfect method would give the conductivity or resistivity of a solution as a function of the concentrations of all solutes present. This is so complicated as to be impractical, and a simpler method must be found which is of acceptable accuracy. The Dunlap method, on the other hand. is too simple because it askmes that at any concentration the contrih-

Jan 1, 1967

By Thomas P. Forbath

THE sudden realization that known sulphur reserves amenable to mining by the Frasch hot water process are nearing exhaustion focused attention on widely scattered surface deposits throughout the world. These deposits are not necessarily of lower sulphur content than ores located underneath Louisiana or Texas salt domes which usually average about 30 pct sulphur disseminated in limestone matrix. Their near surface occurrence, however, renders exploitation by the Frasch process impossible. As is well known, the Frasch process depends on the presence of 500 to 1000 ft of overburden and cap rock above the sulphur deposits to permit melting underground sulphur in place by diffusing hot water under pressures of 200 to 600 psig in the formation and raising the molten sulphur to surface by air lift. This process renders possible the production of pure sulphur which is 99.5 pct pure without any subsequent treatment. Surface deposits contain sulphur in the same range of concentrations as the salt dome deposits, i.e., from 10 to 50 pct sulphur, associated with various gangue materials such as silica, limestone, and gypsum. The pirincipal distinction, then, does not lie in the percentage of sulphur contained in the ore, but in the geological nature of the deposit. A recent study' of the world sulphur supply situation estimated 1950 sulphur production in the free world countries at 5.6 million long tons, of which the United States produced 5.2 million tons, or 93 pct of the total. While America's domestic needs alone would have been covered by national production, about 1.4 million tons were exported during the same year. Despite all the steps which are being taken to restrict use of elemental sulphur and to force the fullest possible development of alternate sulphur sources here and abroad, the deficit in elemental sulphur production will probably increase with time. As a result of intensive prospecting for oil throughout the Gulf Coast area discovery of significant new salt domes is held unlikely. With the growing scarcity of sulphur and what appears to be an inevitable rise in price, recovery from deposits not amenable to Frasch-process mining assumes greater economic importance. Untapped Reserves The most important deposits in this category are located in Sicily, where elemental sulphur occurs in Miocene limestone and gypsum formation. Sulphur content of these ores ranges from 12 to 50 pct with an estimated average of 26 pct. Although quantitative estimate of these reserves is not available it is held that they exceed 50 million tons of sulphur. Similar deposits occur also on the mainland which contribute about one-third of Italy's total current annual production of 230,000 tons, but these are known to be nearing exhaustion. Significant surface deposits of volcanic origin are located in South America, Japan and western United States, silica being characteristic gangue con-stituent. The largest of these deposits are in South America. More than 100 extend over a zone 3000 miles long, paralleling the west coast of South America. 'Total sulphur content of these deposits has been estimated to be as high as 100 million tons. The main islands of Japan also possess at least 40 known volcanic sulphur deposits with probable reserves of 25 to 50 million tons.' Prospected reserves in western United States might amount to 2 million long tons, principal deposits being located in the northwestern part of Wyoming, southern Utah, and eastern California. Volcanic deposits of lesser importance are found around the Mediterranean, in Turkey and Greece, and in Africa, Egypt, Abyssinia, and Somaliland. Beneficiation Methods Different methods of beneficiation have been used in these various locations. In Italy the Calcarone kiln and Gill regenerative furnaces are used exclusively. Both utilize heat liberated by burning part of the sulphur in the ore to liquify or vaporize the remaining sulphur, which is recovered by solidification or condensation. The Calcarone kiln is of conical shape, generally 35 ft in diam at base and 18 ft high. A kiln of 25,000 cu ft capacity burns for about two months and yields about 200 tons of sulphur. The Gill furnace consists of a series of chambers with domed roofs. Sulphur is burned and melted in one chamber at a time and the hot combustion gases are used to preheat the ore charge in the subsequent cell. These furnaces operate on a cycle of 4 to 8 days. The recovery yield of both systems is about 65 pct. Sulphur losses amount to 25 pct through the combustion to sulphur dioxide; about 10 pct is retained in discarded calcines. Ores containing less than 20 pct are not considered suitable as furnace feed. These methods are not only wasteful because of the low recovery obtained, but represent a serious atmospheric pollution problem. Sulphur produced ranges from 96 to 99 pct purity and thus does not match Texas or Louisiana sulphur. Owing to the present shortage, sulphur in the Middle East sells

Jan 1, 1954

By J. A. Rowland, C. W. Funk

THIS investigation is a part of the United States Bureau of Mines work in conserving the Nation's resources. The isothermal sections presented were developed as a guide to a comprehensive investigation of the properties and fabricating characteristics of copper-base alloys containing manganese and tin. These alloys are being investigated as possible replacements for the commercial bronzes containing substantial quantities of tin. Development of the isothermal sections has, therefore, been limited to the area between 0 to 20 pct Mn and 0 to 25 pct Sn. Since Heusler1 reported the ferromagnetic properties of the Cu-Mn-Sn alloys in 1903, the system has been the subject of several investigations. These studies were largely concerned with correlation of magnetic properties and crystal structure.2-6 Vero,' however, established the solidus surface of the system, between 20 pct Mn and 40 pct Sn, by thermal and microscopic methods. The a, ß, and y solid solutions were also observed by Vero and reported to extend into the interior of the ternary system. For the present study, the annotated diagram of the Cu-Sn system prepared by Raynor' and the Cu-Mn diagram by Dean and coworkers9 were used as the binary borders of the ternary system. Procedure The metals used in this investigation were electrolytic manganese, washed cathode copper, and three-star tin. The manganese, produced at the Boulder City pilot plant of the Bureau of Mines, had a purity of 99.9 pct; the copper contained less than 0.004 pct Bi and 0.001 pct S; the tin contained 0.070 pct Pb, 0.070 pct Sb, 0.050 pct Fe, and 0.065 pct Si. Heats of 1 lb were melted in alundum crucibles with a 3-kw induction furnace. The heats were chill-cast in iron or copper molds to form 6-in. ingots 3/4 in. in diameter. Molds were washed with graphite or zirconia, depending on the manganese content, and preheated to 150°C. The maximum impurity content of any alloy was 0.03 pct Fe, 0.02 pct Si, and 0.02 pct Al; in most cases, they contained less than half this quantity of any of these impurities. Composition of the alloys and their treatment prior to homogenizing are shown in Fig. 1. Designations adjacent to each composition refer to the heat numbers. With a few exceptions, ingots containing 20 pct Sn or less were hot-swaged. These ingots were swaged, after a 24-hr preheat at 650°C, to obtain a total reduction of 70 pct in cross section. Intermittent reheating was necessary, and reductions were limited to either 0.025 or 0.050 in. in diameter per pass. Selection of a 200-hr homogenizing treatment was based on extensive experiments which indicated that virtual equilibrium was reached after 150 hr. Structures were retained by quenching in water. Grain sizes in the specimens homogenized at 350" and 450°C were generally small. These homogenizing treatments were preceded by a 50-hr anneal at 650°C to facilitate phase identification by increasing the grain size in these specimens. The specimens were furnace-cooled from this temperature to the homogenizing temperature, and the heat treatment was continued for the standard 200-hr period. The homogenized samples were bisected to provide an internal section for metallographic examination. Optimum definition of the microstructure was obtained by successive polishing and etching. Quarter-strength ASTM copper reagent No. 13 gave the most satisfactory results.10 Filings from interior sections of the samples were used for diffraction studies. These filings were annealed in evacuated glass tubes at homogenizing temperatures for periods exceeding 50 hr and quenched in water. During this treatment the manganese content was reduced by significant amounts, probably by vaporization. This change in composition was considered in applying the diffraction data, which were used only as a means of identifying phases where the data were consistent with the metallographic evidence. Phase Identification The a solid solution is readily distinguished by the copper-colored, twinned, polyhedral grains, as shown in Fig. 2, and a face-centered cubic X-ray pattern. Tukon indentations, using a 25-g load, indicate a Knoop hardness of 155 for a quenched from 650°C. Using white light developed by a Wratten 78 A filter, the ß grains in etched specimens have a dark brown tint in contrast to the lighter copper-colored grains of the a phase. Knoop hardness values of 284 were obtained for the ß structure in specimens quenched from 650°C. The ß structure was also distinguished from the a phase by its body-centered cubic lattice. The acicular structure appearing in Fig. 3 was evident in several alloys of higher tin content after quenching from either 750' or 650°C. Since a trans-

Jan 1, 1954

By K. P. Gupta, A. K. Pal, L. Chandrasekaran, U. P. Singh

The Mn-Sn binary system, investigated at the high-manganese end and between 500° and 1000° C, shows four phases at temperatures below 727"C, namely the u Mn, the p Mn, the Mn3 Sn, and the Mn, Sn phases, while at higher temperatures only the last three phases remain stable. The solubility of tin in a Mn is very small and the maximum solubility of tin in P Mn phase appears to be abollt 10 at. pct Sn. The solubility range of the Mn,Sn and the Mn, Sn phases is 23.0 to 26.0 and 37.0 to 40.0 at. pct Sn, respectively. The lattice parameter of the 13 Mn phase increases with increasing tin content. The Mn, Sn thase is hexagonal with and appears to be basically of the Ni3 Sn type structure except that for quite a few X-ray diffraction lines the calculated and observed relative intensities do not agree well. The Mn-Sn binary system has been studied by several investigators.''~ Their results indicate that three intermediate phases, namely, the Mn3Sn, Mn,Sn, and MnSn, phases, exist at high concentrations of tin. However, so far the proper phase equilibrium has not been established at the manganese-rich end and very little data is available for the composition range between pure manganese and Mn3Sn. Moreover, earlier investigators differ in their opinion about the exact composition range at which the Mn3Sn and Mn,Sn phases appear and some doubt has been cast by some investigators3 regarding the structure of Mn3Sn phase which has been reported to be isotypic with Ni3Sn (Mgscd type) structure. In this investigation an attempt has been made to establish the proper phase equilibrium between pure manganese and Mn + 50 at. pct Sn composition in the temperature range of 500" to 1000°C. PROCEDURE The raw materials used were from three different sources. For exploratory work five alloys containing 5, 10, 15, 20, and 25 at. pct Sn were prepared using 99.9 pct pure manganese and tin supplied by E. Merck & Co., Germany. The rest of the alloys and one more 25 at. pct Sn alloy were prepared using 99.9 pct Mn supplied by Gallard Schlesinger Chemical Mfg. Corp., U.S.A., and 99.999 pct Sn supplied by Semi Elements Inc., U.S.A. Weighed amounts of manganese and tin were melted in recrystallized alumina crucibles in an inert gas (argon) high-frequency induction melting furnace. By careful control of temperature and time of melting the losses were reduced to below 0.2 pct in all cases. Since the losses were very small, no attempt was made here to analyze the samples chemically. Alloys were wrapped in molybdenum foil and sealed in small evacuated fused silica capsules. The alloys were annealed at different temperatures, controlled within + l°C, for sufficiently long periods to attain proper phase equilibrium, and subsequently quenched in cold tap water. The annealing periods used at different temperatures were 15 days at 500°C, 7 days at 600°C, 5 days at 700" and 750°C, 3 days at 800°, 850°, and 900°C, 2 days at 968"C, and two alloys annealed at 968°C were reannealed for 10 hr at 1000° C. From each annealed specimen, a part was utilized for metallographic study while another piece was used for X-ray diffraction study. 1.0 pct HNO, solution and oxalic acid solutions of concentrations 0.05 to 1.0 pct were used for etching Mn-Sn alloys above and below the MnsSn composition, respectively. Since all alloys were brittle, X-ray specimens were prepared using the as-crushed -325 mesh alloy powders. Only one 25 at. pct Sn alloy powder was reannealed in an evacuated silica capsule at 800°C for 5 min and water-quenched. X-ray diffraction patterns . for the Mn3Sn phase with the as-crushed and the reannealed powders did not show appreciable change. Un-filtered iron radiation at 25 kv, 15 ma was used with either Norelco 114.6-mm-diam Debye Scherrer Camera (for phase identification) or Norelco 12-cm-diam symmetrical focusing camera (for lattice parameter determination of the 6 Mn phase). The estimated accuracy of lattice parameter determination for the focusing camera was * 0.001A. RESULTS AND DISCUSSIONS The results of metallographic and X-ray diffraction study made with different alloys are shown in Fig. 1. The variations in lattice parameter with composition for the p Mn and the Mn3Sn phases are given in Tables I and 11, respectively, and the lattice parameter as a function of composition for the p Mn phase is shown in Fig. 2. The lattice parameter of the p Mn phase increases with increasing tin content while for the Mn,Sn phase the data obtained from two two-phase alloys and one single-phase alloy indicate increase in a,, and decrease in c, parameters with increasing tin content. The results, Fig. 1, indicate that the solubility of

Jan 1, 1969

By Mark R. Lamb

THE treatment of slime is of special interest to those engaged in cyaniding gold- and silver-ores. The usual practice is to make as small a percentage of slime as possible. In many instances the slime is given no treatment, but is impounded in dams in the hope that the future will develop some method of economically treating this product. The filter-press was the first step upward from ordinary decantation, but, on account of heavy labor-charge and high cost, its use has been limited to high-grade material. The slime at the Combination mill at Goldfield averages perhaps $20 per ton, and although the values are quickly dissolved the filter-press installation was not entirely satisfactory, resulting in the erection of the canvas-cell filter, developed by Chas. Butters and his staff. This filter is a great improvement over the filter-press, and the following description of it will be of interest to those engaged in treating slime produced in crushing ores, especially in view of the fact that, by the use of this filter, slime can be treated at a lesser cost and with a higher percentage of gold- or silver-extraction than in the ordinary treatment of sand; and, moreover, the initial outlay for an all-sliming plant is less than that for the ordinary sand- and slime-plant. The economy of construction of the Butters filter-plant is clearly shown in Figs. 1 and 2, which illustrate the 20-frame filter of 40 tons capacity, now being installed by the Nevada Goldfield Reduction Co. In this construction the slime-pump is so connected that it can pump to or from either tank or the filter by changing the valve-settings-a combination which is necessary in this particular installation, since sufficient fall is not available for filling and discharging the filter by gravity. Ordinarily, a slime-plant comprises: a filter-box with frames

Jan 1, 1907

By W. T. Burns

Attempts were made in 1908, at the Great Falls Works, to produce ingots direct from the Butte precipitate by smelting the material in a reverberatory refining furnace. The ingots produced in this manner averaged 99.4 per cent. of copper and 0.40 per cent. of arsenic and antimony. The copper content being considered too low, it was decided to cast the copper into anodes and to treat the anodes in the electrolytic plant. No. 1 Furnace, in which the precipitate was treated, was a regular cathode refining furnace previously employed in making wire bar, cake, and ingot from electrolytic cathodes. The hearth of the furnace was 24 by 12 ft., having a capacity of 100,000 lb. of cathodes. The furnace was coal fired, and the copper was dipped by means of trolley ladles. The average composition of the furnace charges was as follows: Per Cent. Peer Cent. Copper. Precipitate,..........7.5.3 40 to 80 Purification cathodes,.....1.9 92 Refinery slag (oxide slag),.. 10.1 50 to 60 Electrolytic cathodes,... 2.1 99.9 Precipitate anode scrap,.... 3.1 99.0 Scrap copper,.........7.5 99.2 100.0 The slag from this furnace amounted to 46 per cent. of the total amount charged and was of the following average composition: Per Cent. Cu..............16.86 SiO2.............34.7 FeO.............18.8 A12O3.............13.0 CaO.............9.4 MgO.............2.7 S............... 0.47 As..............0.11 Sb...............0.006 Ag, oz per ton.....0.18

Jan 1, 1914

By AIME AIME

FRANK T. SISCO is a native of Lawrence, Kansas. He received his earlier education in Clinton, Iowa and his university training at the University of Illinois. His experience in the steel industry has included service in the chemical, testing, and metallurgical departments of the Illinois Steel Co., of South Chicago; the Hess Steel Corp., of Baltimore; the IJ. S. Naval Ordnance Plant at Charleston, W. Va.; and the American Steel and Wire Co., Waukegan, Ill. He severed for seven years as chief of the metallurgical laboratories of the U. S. Army Air Corps at Wright Field, Dayton, before coming to New York in 1930 to become metallurgist and editor of the Alloys of Iron Research, the position he now holds.

Jan 1, 1940

(Under this heading will be published notes sent to the Secretary of the Institute by members or other persons. ) The governing body of the Imperial College of Technology, London, S. W. (Royal School of Mines), will shortly proceed to the appointment of a Professor of Metallurgy. This important Chair has been filled in the past by such eminent metallurgists as John Percy, Sir William C. Roberts-Austen, and others. A large manufacturing company having two iron foundries, blast furnace, etc., requires the services of a chemist, metallographist, and metallurgical engineer qualified to give advice on cast iron practice, and to prescribe remedies for difficulties often met with in the usual iron foundry. The position of converter blower at a side-blow Bessemer steel casting plant is vacant.

Jan 5, 1913

On October 20, 1964, in his Waldorf Towers suite, death came to Herbert Clark Hoover, 31st President of the United States. At his bedside were his two sons, Allan A. and Herbert Hoover, Jr. Many years of humanitarian and public service earned him the gratitude of millions here and abroad. The Nation mourned his death. For the mining fraternity and AIME in particular, the life of Herbert Hoover held particular meaning. During the years from 1895 to 1914 when he was actively engaged in mining, he conducted himself with extraordinary competence, energy and integrity. In the years following, when his qualities of administration and leadership literally forced him into the limelight, he continued his interest and lent his talents for the benefit of the engineering profession.

Jan 11, 1964

By R. M. Schuster, F. M. Benavides

A plan is prepared to develop and operate a four mine complex over a 30-year period. The mining complex is in the vicinity of two existing processing plants and the most economical means of transporting the raw minerals to either or both of these processing plants is sought. The mines and processing plants are all within a radius of 5.6 km (3.5 miles). Existing transportation facilities consist of county and state roads and federal highways as well as haul roads in some areas. A navigable river meanders through the area under study and there are two railroad systems servicing the area. The analysis shows the economic and operational feasibility of using an overland conveyor in lieu of other means of transportation.

Jan 1, 1983

By Theodore H. Eyde

The Superior Oil Co. drilled two holes, 4720 and 5940 ft, respectively, using conventional oilfield rotary drilling equipment. The results indicated that large rotary drills can be adapted to mineral exploration purposes with few modifications. The principal advantages are: a large hole diameter, excellent samples of cuttings, a geolograph which is a constant drilling record, rapid penetration even in hard rocks, and the ability to use electric logs for measuring the properties of the rock being penetrated. The disadvantages are the high cost of mobilization, drilling, and oilfield service at remote locations. Though the oilfield sampling techniques are adequate for identifying lithology, improvements are necessary to obtain representative samples for analytical purposes. As deeper targets are explored, mineral exploration organizations will have to resort to rotary drilling.

Jan 1, 1975

By AIME AIME

PROBABLY no man has been of greater service to the Institute and has kept more in the background than Henry Krumb. A Vice-President continuously) for the last eleven years, apparently neither his picture nor a biographical sketch ever have adorned these pages and were he forewarned in the present instance he would order us to "forget it." He is a Columbia School of Mines man, Class of '98. He worked underground at Rossland, B. C. for a time, then for a year and a half a. chief engineer of the famous Camp Bird at Ouray, Colo. For three wars he was examining engineer for the Guggenheims and since 1901 has been an independent consulting engineer with experience throughout the Americas.

Jan 1, 1939

THE metal iron has physical arid chemical properties which are some- what different from those of steels, but a knowledge of the pure metal is a useful starting point in studying the behavior of steels. Solid iron exists in two different crystalline forms, called alpha (a) and gamma (?) iron the former being the ordinary form at room temperature. When heated above the "A2 point," 769 C (1416 F), the metal becomes nonmagnetic, and at the "A3 point," 911 C (1672 F), it undergoes a change in crystalline form to become gamma iron. 111 the range 769 to 911 C, it is sometimes called beta (ß) iron to distinguish it from the magnetic alpha, but the two are identical in crystal structure. When gamma iron is further heated to 1392 C (2538 F), it changes into delta (d) iron, which is identical with beta or nonmagnetic alpha iron. Pure iron melts (in an inert atmosphere) at 1536 C (2797 F); the melting point is generally lowered by alloys or impurities. PHYSICAL AND THERMAL PROPERTIES OF IRON Many of the properties of iron undergo abrupt changes at the transition temperatures.1 For example, in heating through the A3 point, a sudden shrinkage in volume occurs. Similarly, melting is attended by sharp physical changes such as an increase in volume. Table 16-1 contains a summary of the measured physical properties of pure iron at several temperatures. Usually the properties change smoothly with the range of stability of a given form, but this is not always true. For ex- ample, the specific heat reaches a sharp maximum at the A2 point al- though no transformation occurs. (See Fig. 14-1.) Solvent Powers of Liquid Iron. Liquid iron is capable of dissolving most of the other metals and many nonmetallic elements as well. Of the metals that are liquid at the melting point of iron, only lead, silver, and bismuth refuse to dissolve to an appreciable extent. Table 16-2 contains a list of the common elements classified with respect to their solubility in liquid iron. Those which are very soluble or practically insoluble require no further discussion, but several of the nonmetals that are gases and exhibit limited solubility are of considerable importance to us and merit more detailed consideration.

Jan 1, 1964

By Nicholas J. Grant, Noboru Komatsu

Electron-micrographic studies and hardness tests were made of an internally oxidized dispersion-strengthened Cu-12 vol pct SiO2 alloy follozuing longtime annealing at elevated temperatures. The rate of coarsening of SiO2 particles and the relationship between hardness and interpavticle spacing were Dispersion -strengthened alloys exhibit a remarkable stability of structure and properties at elevated temperatures compared with conventional alloys; nevertheless, when they are exposed to high measured. The rate of coarsening of SiO2 particles is controlled by the decomposition of SiO, rather than by the diffusion rates of silicon or oxygen through the copper matrix. The mechanism of coarsening was also studied by using a diffusion-couple technique. temperatures, weakening takes place as a result of coarsening of the dispersed phase. The theoretical treatment of the coarsening of dispersed phases has been treated by Greenwood,1 wagner,, and others3 by considering the interfacial energy of dispersions and the diffusion of solute atoms. Recently, experimental work on the coarsening of relatively insoluble, finely dispersed cobalt particles in copper was reported by Livingston,4 and on aluminum oxide in nickel by Dromsky, Lenel, and Ansell;5 according to these investigations, the rate

Jan 1, 1964

By F. P. Haver, M. M. Wong

When a mixture of chalcopyrite concentrate and lime is heated in air, the following reaction takes place: 2CuFeS2 + 4Ca (OH)2 + 8 ½ 02 ; 2Cu0 + Fe203 + 4CaS04 + 4H20. Ca (OH)2 is the only common reagent that will adequately control SO2 evolution in roasting. Other additives tried, including CaO, CaCO3, Na2CO3, NaOH, MgO, Mg (OH)2, and MgCO3, were not effective. All reacted with sulfur, but not to the extent that will be required to meet the new control regulations for sulfur oxide emissions. For example: when the concentrate was heated with the theoretical amount of limestone (CaCO3), about 55% of the sulfur was retained in the calcine; whereas use of the stoichiometric amount of Ca (OH), converted approximately 99% of the sulfur to CaSO1. In most of the experiments, Ca(OH)2 was formed by adding CaO to concentrates containing enough water to simulate filter cake from flotation circuit. This eliminated the need for drying the concentrate and utilized the heat of hydration of CaO.

Jan 6, 1972

As a service to its readers, MINING ENGINEERING is publishing the mineral industry portion of the 1957 list of Accredited Undergraduate Engineering Curricula in the United States. This list, revised annually, is issued by the Education Committee of the Engineers Council for Professional Development, and covers all phases of engineering education. The AIME is actively represented on ECPD, which was jointly organized by the Founder Societies and various other groups interested in the professional recognition of engineers. ECPD is merely authorized by its constituent organizations to publish a list of accredited engineering curricula for use as desired by those agencies which require such a list. It has no authority to impose any restrictions or standardizations on schools. (Date following school refers to year of initial accrediting.)

Jan 7, 1958

PLANS for development of the San Manuel ore- body have been worked out and are progressing parallel with construction of the surface plant. This property lies in the Old Hat mining district, Pinal County, Ariz., about a mile south of Tiger where the St. Anthony mine once produced gold, molybdenum, vanadidn, lead, and zinc. Two development shafts have been completed and serve as ingress for preparation of the 1475 haulage level and 1415 grizzly level. Three other shafts are being put down. The Nos. 3A and 3B shafts will provide the ore hoisting facilities. The No. 4 shaft, now being sunk, will be at the main mine service and supply yard, with primary function to provide the mine with supplies.

Jan 7, 1954