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By Alan Morris

IN a previous paper1 the results of cutting tests on free-cutting brass rod were reported. Investigation was made of the effects of variation in lead content, microstructure and cold drawing. The author's reply to the discussion of that paper includes a study of the directional cutting properties of drawn and annealed rod. The present paper records the results of tests made on series of samples in which the tin, iron and copper contents have been varied. The apparatus, used and method of testing are described in detail in the previous paper. The machine is of the pendulum-operated milling type used and described by Airey and Oxford.2 Angular readings of the height to which the pendulum rises after falling from a fixed position, both when a cut is made and when the pendulum swings freely, give the basis for calculation of the energy consumed in making a cut. Careful weighing of the sample before and after cutting gives a measure of the amount of metal removed. The results are expressed in foot-pounds per cubic inch of metal removed, and have been called "unit cutting energy." In the course of the discussion of the author's earlier paper, W. B. Price reported large-scale tests involving some 50,000 lb. of material. These tests indicated that the addition of as much as 0.6 per cent of iron and 0.25 per cent of tin had apparently no adverse effect on the high-speed cutting characteristics of the rod. The rod also contained a small amount of nickel. The test results reported here tend to confirm his observations.

Jan 1, 1933

By Paul M. Tyler

YEAR after year, the kaolin industry of the United States has been setting new production records and making better products. High-grade paper, pottery, and rubber clays are produced in this country mostly in the South. Georgia alone contributes over 70 pct and South Carolina almost 20 pct of the total domestic output. Residual kaolin is mined in North Carolina, highly plastic but naturally sandy Tertiary (Eocene) potting clays are worked in north central Florida, and good white clays are produced in several other states, but the main sources of kaolin or china clay have been numerous deposits in the Tuscaloosa (Upper Cretaceous) formation. This formation of generally sandy sediments is called the Middendorf member in older geologic reports and corresponds in age with some of the New Jersey clays. As shown in fig. 1, it crops out almost continuously in a generally southwesterly direction across South Carolina and Georgia and into Alabama. Clay is mined from this formation in all three states but the principal producing centers lie within about 10 miles of a straight line drawn between Aiken, S. C., and a point about 10 miles south of Macon, Ga. The white kaolins of the South were recognized and used prior to the Civil War but suitable treatment processes were not introduced until World War I when imports, chiefly from England, were curtailed. Although imports of high-grade clays were resumed after 1918, the domestic industry managed to treble its prewar production record during the early 1920's and has continued to grow. Whereas the 1909 to 1913 average total production in the United States was only 132,104 short tons valued at $705,352 f.o.b. mines, the output in 1948 was 1,568,848 tons worth $19,756,738.

Jan 6, 1950

By Willard L. Hunter

Four solvent extraction systems were studied to determine their efficiency jor extraction and separation of tantalum and columbium. Aqueous feed solutions of varying HF-HCl concentrations and metal content were contacted with equal volumes of cyclohexanone, 3-methyl-2-butanone, and 2-pentanone and solutions of varying HF-H2S04 concentrations were contacted with equal volumes of 2-pentanone. One multistage continuous test was made in a polyethylene pulse column using cy clohexanone as the organic phase. In each system studied, columbium and tantalum purities in excess of 95 pct with respect to each other were obtained in single-stage tests at low acidities in the feed solution. Separation factors ranging from 1700 to 2400 were obtained when rising HF-HCl mixtures in the aqueous phase. Best results were obtained when a solution of HF-H2S04 was used as the aqueous phase and 2-pentanone as the organic phase. A separation factor in excess of 6000 was obtained in one stage with aqueous solution concentrations of 2 _N HF and 2N H2S0,. When acid concentrations were increaszd to 52 HF and 10 _N H2S0,, 99.9 pct of the tantalum and 98.2 pct of the columbium initially present in the feed solution were transferred to the organic phase. The separation of columnbium and tantalum obtainable by means of the solvent extraction systems presented in this paper was found to corn -pare favorably with other systems, including the HF-H2SO4-methyl isobutyl ketone system currently used by most producers for the extraction and separation of these metals. TANTALUM and columbium are always found together in minerals of commercial significance, although the proportion of the two metals in ores varies within broad limits. Columbium is estimated to be 13 times more abundant than tantalum. Five methods generally employed for the separat:ion of these metals are: 1) fractional crystallization (the Marignac process),2 2) solvent separation, 3) fractional distillation of their chlorides, 4) ion exchange, and 5) selective reduction. Of these methods, the one currently used by industry to the greatest extent is that of solvent separation. One of the early technical developments in solvent separation of tantalum from columbium was reported by the Bureau of Mines: the HF-HC1-methyl isobutyl ketone system; data were presented for both laboratory and pilot-plant experimentation.3 Of twenty-eight organic solvents tested for their ability to extract tantalum from an HF-HC1 solution of columbium and tantalum, 3-pentanone (diethyl ke-tone), cyclohexanone, 2-pentanone, and 3-methyl-2-butanone were chosen for further study. Data on the HF-HC1-diethyl ketone system has been published4 and data describing the use of cy clohexanone, 2-pentanone, and 3-methyl-2-butanone as the organic phase are included in this report. RAW MATERIAL The source of tantalum and columbium oxides for this study was ('Geomines" tin slag from the Manono Smelter, Cie Geomines, Gelges, S.A., Congo. In order to extract the valuable Ta-Cb content, the slags were carbided, chlorinated, and the sublimate from chlo-rination was hydrolized and washed free of chloride with water. The washed material was air-dried and stored in a stoppered container. Throughout the paper, "feed material" refers to this mixture of hydrated oxides which was employed because of its high solubility in aqueous solutions. Typical analysis of the hydrated oxides is shown in Table I. I) HF-HC1-CYCLOHEXANONE SYSTEM Batch Separation. Effect of Acid Concentration. To determine the effect of varying the acid concentration upon the transfer of tantalum and columbium, a series of tests was made in which approximately 2.5 g of feed material was added to 25 ml solutions of 2, 4, 6, 8, and 10 N HF and 0 through 5 N HC1. Tantalum pentoxide concentration of the solu%ons was approximately 21 g per liter and columbium pentoxide was 14 g per liter. These starting solutions were shaken with equal volumes of cyclohexanone in 100 ml polyethylene bottles for 30 min. The phases were carefully separated in 125 ml glass separatory funnels. The time of contact of the solutions with the separatory funnels was kept at a minimum to reduce silica contamination. The measured phases were separated into 400 ml polyethylene beakers and the metal contents of each were precipitated by addition of an excess of ammonium hydroxide. Precipitate from each phase was filtered on ashless filter paper, ignited at 800" to 1000°C for 45 min, weighed, and analyzed by X-ray fluorescence.5 Data tabulated in Table I1 and illustrated in Fig. 1, show that maximum separation of tantalum from columbium for each HF concentration was obtained with no HCl present. The purest tantalum product was obtained with some HCl present. The highest separation factor was obtained at 2 N HF and

Jan 1, 1970

By D. A. Elgillani, M. C. Fuerstenau

In this paper, oxidation potentials measured in the presence of various concentrations of cyanide, ferro-cyanide, and ferricyanide and ethyl xanthate at various values of pH are related to flotation response. Eh-pH diagrams are presented to show that the formation of surface ferric ferrocyanide is probably responsible for depression when cyanide is added. The influence of cyanide on the depression of pyrite with xanthates as collector has been the subject of a number of investigations,'-6 and several theories on the mechanism of depression have evolved from these studies. Wark and Cox7 and Gaudin8 have suggested that the depressing effect is due to a competition of cyanide ion with xanthate ion for the surface. Cook and his colleagues9-11 have explained this phenomenon in terms of competition between hydrocyanic acid and xanthic acid. Sutherland 12 has shown that although both of these theories accurately describe the relation between pH value and cyanide addition at constant collector addition, they fail to describe the relation between pH value and the amount of collector required to cause flotation. Taggart 13 suggested that depression in these systems is due to the formation of a reaction product between ferric ion at the pyrite surface and ferrocyanide ion derived from solution. Majumdar4,6 has attempted to prove this hypothesis by measuring the contact angles of pyrite in the presence of 25 mg per liter ethyl xanthate and different concentrations of potassium ferrocyanide and ferricyanide. In all cases the contact angles were quite high up to pH 10. These results indicate that pyrite should not be depressed by either potassium ferrocyanide or ferricyanide. In view of these facts, Majumdar has assumed that the compound Fe(CN)2 forms at the surface. Gründer and Bornl4 have stated that depression may be due to the formation of the compound K2Fe(II)Fe(CN)6 at the pyrite-solution interface. This compound is thought to be an interaction product between the K2Fe(CN)6-2 ion from solution and the Fe++ ion at the pyrite surface and, accordingly, K4Fe(CN)6 should depress pyrite at least as effectively as KCN. This was proven experimentally, but there was no simple relation between the depression of pyrite and the concentration of either KCN or K4Fe(CN)6 in solution. In view of the many mechanisms that have been proposed for pyrite depression by cyanide, it is apparent that a clear understanding of the phenomena occurring in these systems is lacking. One reason for this may be the fact that the species responsible for pyrite flotation in the presence of xanthate is not the xan-thate ion but rather dixanthogen.15 Since the oxidation of xanthate to dixanthogen is dependent on the oxidation potential of the solution, it would seem that knowledge of these potentials would be a requisite to understanding the pyrite-xanthate-cyanide system. It is the object of this paper to measure both the oxidation potential and pH of the pyrite systems in the presence of various concentrations of cyanide, ferrocyanide, and ferricyanide and xanthate and to relate these values to flotation response. EXPERIMENTAL MATERIALS AND 'TECHNIQUES In the experiments discussed here, pure potassium ethyl xanthate was used as collector, and reagent grade potassium cyanide, potassium ferrocyanide, and potassium ferricyanide were used as depressants. Reagent grade HC1 and KOH were added for pH adjustment. Conductivity water, made by passing distilled water through an ion exchange column, was used in all experimental work. Two natural samples of pyrite were used in the investigation. Sample preparation for flotation included dry grinding with a mortar and pestle and sizing the product to 100 x 200 mesh. Prior to flotation, a 0.75-gm sample of pyrite was added to a solution containing a known amount of depressant at the desired pH value, and the system was conditioned for 4 min. Following this, a known amount of collector was added and the system was conditioned for another 4 min. The pH — termed flotation pH - was measured; the pulp was transferred to a Hallimond cell, and flo-

Jan 1, 1969

By C. A. Stickels

A heat of electrolytic iron, to whzch alunzinutn and nitrogen had been added, was hot-rolled, cold-rolled 90 pct, and recrystallized at temperatures from 500" to 700°C. Primary recrystallization textures appear to arise from competitive growth of two types of nuclei: 1) those having orientations belonging to the "usual" primary recrystallization texture found in riming steel, and 2) those with the {111} (110) ovientation. Development of a (111}(1 10) component in the primary recrystallization texture occurs only over a certain interval of isothermal recrystallizatzon temperatures when the material is supersaturated with respect to the precipitation of AlN. Lowering the degree of supersaturation depresses the temperature interval in which a (111)(110) component occurs. An elongated, 'pancake-shaped" recrystallized pain structure and a marked delay in the start of recrystallization were found in all specimens which were supersaturated with respect to A1N precipitation after cold work, regardless of their recrystallization texture. ONE of the consequences of killing low-carbon steel with aluminum is a significant change in recrystallization behavior. About 15 years ago, Solter and eatttiel showed that this behavior was largely controlled by aluminum and nitrogen in the steel. If complete precipitation of A1N was prevented before cold rolling, an increased "recrystallization temperature" was observed in subsequent. annealing, and the recrystal-lized grains were not equiaxed. Leslie et a1.2 studied this phenomenon in some detail and clearly demonstrated the relationship between A1N precipitation, recrystallization kinetics, and the development of "pancake-shaped" grains. It has also been known for some time that aluminum-killed steels, processed to produce elongated "pancake" grains, develop a (11 I}( 110) primary recrystallization texture. This texture has not been found in iron or low-carbon rimming steel as a primary texture4j5 but has been observed following grain growth in electrolytic iron.5 The present work was undertaken to study in more detail the effect of A1N supersaturation on recrystallization textures in iron. LITERATURE REVIEW The deformation texture in heavily rolled iron has been studied in detail by Bennewitz.~ The texture consists primarily of a partial fiber texture about a (110) axis in the rolling direction, designated here as fiber texture A. It includes the range of orienta- tions (111)[110] - (001)[ 110] - (11l)[110]. A weak secondary texture also is present.6 This is a duplex partial fiber texture about two (110) fiber axes located 60 deg from the rolling direction and 30 deg from the sheet normal. The range of this texture, designated here as fiber texture B, about the [101} fiber axis is (112)[110] - near (545)[252] - (211:1[011] *The range given here follows Bennewit~.~ A few pole figures from re-crystallized material indicate a broader range than this.' However, the components which are strongest in the recrystallization texture are in this range.'________________________________________________________ Primary recrystallization textures in unkilled steels can be accounted for by growth of members of fiber texture B present in the deformed metal.5 However, while members of fiber texture B dominate the primary texture, other orientations survive primary recrystallization as well. In particular, some {111}(110) members of fiber texture A must also grow during primary recrystallization, because a well-defined {1ll)( 110) texture develops during subsequent grain growth at 700°C.5 The unusual recrystallization behavior of deformed supersaturated solid solutions has been attributed to: 1) retention of the solute in solution,' 2) formation of coherent, preprecipitation solute clusters prior to and during re~r~stallization,~ and 3) formation of a precipitate prior to and concurrent with recrystallization.'~-'~ When aluminum is supersaturated with iron, the difference in grain boundary mobility between general high-angle boundaries and certain special coincidence site boundaries is apparently eliminated.' In aluminum-killed steels, precipitation of A1N can take place at ordinary subcritical recrystallization temperatures. The rate of precipitation increases with increasing aluminum or nitrogen contents.2'13 There is some doubt, however, as to whether true precipitates form during the time at temperature needed to complete recrystallization. Leslie ef a1.2 found that precipitation in one steel was complete after about 100 min at 700GC, or after about 1000 min at 650GC, as measured by chemical analysis for AlN. Aoki et a1.,13 using internal friction for dissolved nitrogen, showed that a large fraction of the dissolved nitrogen was removed from solution within a few minutes annealing time at temperatures from 400" to 800°C. However , the rate of formation of AlN, as detected bv chemical analvsis. was much slower than the apparent rate of nitrogen removal. Hasebe,'~~ using carbon extraction replicas, has identified A1N precipitates by electron diffraction in a 0.2 C steel, solution-treated at 1300°C and annealed 2 hr at 700°C. Borchers and kim,I6 also using a replication technique, observed precipitates after annealing treatments as short as 2 min at 640°C. However, Leslie et a1.' state that no A1N precipitate can be seen while recrystallization is being inhibited in aluminum-killed steel.

Jan 1, 1967

By J. F. Smith, J. E. Davison

The heat capacity of CaMg2 was measured over the temperature interval, 4.8° to 287°K, by the technique of low-temperature adiabatic calorimetry. Heat content measurements were performed with a drop calorimeter over the temperature interval, 273" to 673°K. From these data the thermodynamic functions, (FT - H0)/T, ST - So, and & - Ho, were evaluated. A third-Law calculation of the standard entropy of formation of CaMg2 yields a value of -0.25 * 0.06 cal per (°K g-atom) , and the free-energy function derived from this study when combined with existing equilibria data yields a value for the standard enthalpy of formation which is in agreement with direct calorimetric enthalpy measurements. The accompanying paper' shows that the enthalpy of formation of CaMg2 has been determined with good precision by three different calorimetric techniques.'-= TWO independent determinations of the Gibbs free energy of formation of CaMg2 have also been made; both determinations were based on vapor pressure measurements, being in one case hydrogen vapor pressures over ternary Ca-Mg-H alloys4 and in the other case magnesium vapor pressures over binary Ca-Mg alloys.5 The present determination of heat capacity of CaMg2 below room temperature and of the heat content of CaMg2 above room temperature was undertaken to provide supplementary data. These data are useful in their own right but can in addition be used to evaluate an entropy of formation for CaMg2 which, because of the interrelation of free energy, enthalpy, and entropy, can be used as a check of the self-consistency of the composite of the presently available information. LOW-TEMPERATURE HEAT CAPACITY The heat capacity of CaMg2 was measured over the temperature interval 4.87° to 286.64°K in an adiabatic calorimeter. The physical details of the calorimeter and the experimental procedure for measuring the heat capacity of a specimen have been adequately described by Gerstein et a1.6 The source and purity of the calcium and magnesium are described together with the methods of sample preparation and chemical analyses in the accompanying paper.' Results of chemical analyses of the material which was used in the present investigation are shown in Table I. These analyses show that, on the basis of the published phase diagram,7 the heat capacity sample contained a slight excess of a calcium while the heat content sample contained a slight excess of magnesium. However, in both cases the excess was small, and X-ray diffraction patterns showed reflections which were without exception attributable to CaMg2. The sample which was used for heat capacity measurements weighed 69 g while the sample container and addenda weighed 132 g. The sample was in the form of annealed powder, 50 to 60 mesh, and was sealed into the sample container under 0.1 atm of helium. Copper fins inside the sample container facilitated thermal equilibrium of the powdered Sample. Time intervals of the order of 10 min were required for thermal equilibration, and such times are normal for this calorimeter regardless of the form of the sample. The observed heat capacities were corrected for the small excess of a calcium through use of the heat capacity values tabulated by Hultgren et a1.8 The corrected heat capacities are tabulated as a function of temperature in Table II. The free-energy function and the absolute entropy of CaMg2, which were calculated from the experimental heat capacity data, are listed in Table 111. A smooth curve was fitted to a plot of the experimental values of the heat capacity and in only two instances above 30°K did the plotted points deviate from the curve by more than 0.2 pct. Below 10°K the deviation of several of the points was as much as 50 pct. These large percentage deviations were attributed to the small value of the heat capacity and to the low sensitivity of the platinum resistance thermometer in this temperature range. The deviations in the region of 10°to 30°K were less than 5 pct. Although the percentage deviations of some of the low-temperature measurements are large, the actual value of these deviations is small since the magnitude of the heat capacity in that temperature range is small. The error in the value of the third-law entropy at 298.15°K was estimated to be less than 0.01 cal per (°K g-atom). A value of -0.25 ±0.06 cal per (°K g-atom) was obtained for the standard entropy of formation at 298.15°K from the relation:

Jan 1, 1969

By H. W. Imholz

Active interest in the North Central Texas area centered in the development of the Palo Pinto limestone-producing zone, near the town of Avoca, in the northeast part of Jones County. This producing horizon was discovered in 1937, in the Iron Mountain Oil Company's well No. 1 Olander, in sec. 196, B.B.B. & C. survey, at 3238 to 3250 ft. During the development of this pool, it was generally referred to as the Avoca pool. As a second pool has been discovered in the same producing zone, also near the town of Avoca, the writer has taken the liberty of changing the name of the original Avoca pool to Avoca-Olander pool. Forty producers had been drilled in the Avoca-Olander pool by the end of 1938; their initial ranged from 28 to 300 bbl. per hour. Stimulus was given to the interest in the development of the Palo Pinto limestone when the Ungren-Frazier No. 1 Griffin, in sec. 199, B.B.B. & C. survey, Jones County, found the limestone from 3225 to 3250 ft., with an initial production of 931 bbl. per day. (This new pool has been given the name of Avoca-Griffin pool, by the writer.) The best well to be completed, up to the present time, is the Bert Fields No. 1 Taylor in sec. 199, B.B.B. & C. survey, which had an initial production of 748 bbl. in 3 hours. The Owen-Snebold No. 1 McKeever in sec. 158, B.B.B. & C. survey, in northwest Shackelford County, added to the interest in the Palo Pinto limestone when it extended the Ivy pool with a well in the Palo Pinto limestone from 3145 to 3183 ft., and was estimated good for 6,000,000 cu. ft. of gas and 6000 bbl. of oil per day. During 1938, Stonewall County obtained its first commercial producer when the Stonewall Oil Co. in No. 1 Parriott discovered the Palo Pinto limestone saturated from 5169 to 5176 ft. and tested 371 bbl. per day. Two new producing horizons were discovered about 2 miles east of the old Noddle Creek pool, Jones County; the first, the Humble Oil & Refining Co. No. 1 Irvin, in sec. 48, block 18, T. & P. R. R. Co. survey, which had saturated limestone from 2550 to 2567 ft. and was good for 100 bbl. per day; the second, in the Humble No. 1 Horton, sec. 40, which

Jan 1, 1939

By H. W. Imholz

Active interest in the North Central Texas area centered in the development of the Palo Pinto limestone-producing zone, near the town of Avoca, in the northeast part of Jones County. This producing horizon was discovered in 1937, in the Iron Mountain Oil Company's well No. 1 Olander, in sec. 196, B.B.B. & C. survey, at 3238 to 3250 ft. During the development of this pool, it was generally referred to as the Avoca pool. As a second pool has been discovered in the same producing zone, also near the town of Avoca, the writer has taken the liberty of changing the name of the original Avoca pool to Avoca-Olander pool. Forty producers had been drilled in the Avoca-Olander pool by the end of 1938; their initial ranged from 28 to 300 bbl. per hour. Stimulus was given to the interest in the development of the Palo Pinto limestone when the Ungren-Frazier No. 1 Griffin, in sec. 199, B.B.B. & C. survey, Jones County, found the limestone from 3225 to 3250 ft., with an initial production of 931 bbl. per day. (This new pool has been given the name of Avoca-Griffin pool, by the writer.) The best well to be completed, up to the present time, is the Bert Fields No. 1 Taylor in sec. 199, B.B.B. & C. survey, which had an initial production of 748 bbl. in 3 hours. The Owen-Snebold No. 1 McKeever in sec. 158, B.B.B. & C. survey, in northwest Shackelford County, added to the interest in the Palo Pinto limestone when it extended the Ivy pool with a well in the Palo Pinto limestone from 3145 to 3183 ft., and was estimated good for 6,000,000 cu. ft. of gas and 6000 bbl. of oil per day. During 1938, Stonewall County obtained its first commercial producer when the Stonewall Oil Co. in No. 1 Parriott discovered the Palo Pinto limestone saturated from 5169 to 5176 ft. and tested 371 bbl. per day. Two new producing horizons were discovered about 2 miles east of the old Noddle Creek pool, Jones County; the first, the Humble Oil & Refining Co. No. 1 Irvin, in sec. 48, block 18, T. & P. R. R. Co. survey, which had saturated limestone from 2550 to 2567 ft. and was good for 100 bbl. per day; the second, in the Humble No. 1 Horton, sec. 40, which

Jan 1, 1939

By Joseph W. Richards

(New York Meeting, February, 1912.) THE Ashio copper-mine of the Furukawa Mining Co. is situated 18 miles from Nikko, and 109 miles north of Tokyo, near the center of Japan. The mine-waters are run over scrap-iron, whereby most of the copper is precipitated as cement copper, leaving iron sulphate with some copper sulphate in solution. After using in the wet concentrating ore-dressing plant, the waste water contains clay and ore-slimes. The only avenue of disposal of this water is the Watarese river; but as this flows through valleys in which the water is used for irrigating the rice-crops, it is necessary to purify and clarify the water before discharge. The water amounts to from 600 to 700 cu. ft. (from 17 to 20 cu. m.) per minute, and averages about 0.00025 per cent. of copper, representing more than 18 tons of copper per month. The water carries H2SO4, FeSO4, Fe, (SO,),, CuSO4 and other metallic sulphates in solution, and clay, ore-fines, slimes and basic. iron sulphates as suspended matters. To purify and clarify the water, precipitation by milk of lime has been used, in three different ways, as follows: I. First Method, Copper not Recovered. Milk of lime was run into the water, and the mixture run through six settling-ponds, and then through a sand filter into the river. Fig. 1 is a view of the settling- and filtering-ponds. In the first pond, sand and metallic hydrates free from copper collected; in the next five ponds, slime, calcium sulphate, and metallic hydrates carrying copper settled out. The effluent carried from 0.1 to 0.05 mg. of copper per liter

Jun 1, 1912

By C. D. Rubert, W. J. Parton

A flotation plant was placed in operation in 1945 at the Tamaqua colliery of the Lehigh Navigation Coal Co. Inc., to recover fine anthracite which is discharged from the cleaning plant with the washery water.‡ In addition to the recovery plant, facilities were provided for the complete clarification of the washery water, recirculation of this clarified water for breaker use, and disposal of the refuse solids in a waste impounding dam. Fig I shows a view of the settling tanks and the flotation plant. Flow Diagram The flow diagram for the complete recovery project is given in Fig 2. All the waste water from the cleaning plant is collected in a sump A outside the breaker structure, from which a 7000 gpm pump B delivers it through a 16 in. line to a 75-ft diam hydroseparator C. The hydroseparator, acting as a classifier, is set to give a separation at approximately 200-mesh. The minus 200-mesh overflow is discharged by gravity into the 180-ft thickener D for final settling of the solids. The classifier underflow sludge is pumped to the flotation plant by a 6-in. quadruplex diaphragm pump E. A 10-ft diam hydroseparator F is used to make a second classification on the sludge solids at approximately 28-mesh. The underflow from this classifier containing the coarse particles is elevated by a 6-in. diaphragm pump G and is discharged on two 4- by 6-ft vibrating screens II. These screens are dressed with approximately 28-mesh cloth. Screen oversize product flows directly to the centrifuge that dewaters the cleaned froth. The screen underflow and the overflow from the 10-ft classifier flow to the flotation circuit. The flotation feed is conditioned in a 12 by 12-ft agitator tank J at a pulp density of I:2.33 (30 pct solids). Reagents are fed to the circuit by reagent feeder I. The conditioned pulp is then diluted to a pulp density of approximately I: 5 (18 pct solids) and is fed to two banks of four 100-cu ft froth flotation cells K for primary coal recovery and cleaning. Additional reagents from reagent distributor L are added to the feed weir-box of Nos. 2, 3, and 4 cells of each bank of flotation machines. Clean coal removed from these cells flows directly to the clean-coal sump. Dilute tailings from the primary cells is re-thickened in a 30-ft diam thickener M and is then retreated in a scavenger circuit for final recovery of coal. The scavenger circuit consists of a 6- by 6-ft agitator tank N and two banks of four 24-cu ft froth flotation cells 0. Coal froth from these cells flows to the dewatering equipment and the tailings flows to a vertical refuse pump which delivers it to the 180-ft diam waste thickener D. Dewatering of the coal products is accomplished in a 54- by 72-in Bird centrifugal filter P. The cake product is dropped onto a conveyor and is loaded

Jan 1, 1949

By Irving Cadoff, Kurt Komarek, Edward Miller

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

Jan 1, 1961

By E. Miller, J. M. Eldridge, K. L. Komarek

The thermodynamic properties of- liquid Mg-Pb allojs have been determined by an isopiestic method from 18 to 42 at. pct Pb and by a thevmodynumic analysis of the phase diagram frofiz 0 to 100 at. pct Pb. In the overlapping composition range the magneszupn activity values derived by the two methods are in excellent agreement, thus mutually supporting the correctness of the Mg-Pb phase diagram previously determined by the authors and of the acticitzes deriz~ed isopiestically. SEVERAL studies have been made of the thermody-namic properties of liquid Mg-Pb alloys. Wagner and Engelhardt,1 Lantratov,2 Knappwost,3 Sryvalin et al.* and Wilson5 used the electromotive-force method, Scheil and Wolf8 and Schmahl and Sieben" applied the transpiration technique, and the present authors8 utilized the phase diagram to determine the activity of magnesium in the Mg-Pb system. The results of these studies are not in very good agreement and could not be incorporated into a general systematic analysis of the thermodynamic properties of the homologous series of the Mg-Group IV B systems Mg-Pb,8 Mg-Sn,° Mg-Ge,10and Mg-Si.11 The thermodynamic properties of the Mg-Pb system were therefore re investigated by an isopiestic technique which has been developed for similar studies in the Mg-SnB and Mg-Ge10 systems- Lead specimens, contained in iron or graphite crucibles, were heated in a temperature gradient in an evacuated and sealed titanium reaction tube and equilibrated with magnesium vapor of known vapor pressure to form alloys of compositions which depended on their temperatures, the temperature of the magnesium vapor source, and the thermodynamic properties of the system. EXPERIMENTAL PROCEDURE Magnesium (Dominion Magnesium Ltd., Toronto, Canada) of 99.99+ pct purity and lead (American Smelting and Refining Co.) of 99.999 pct purity were used. In the early experiments, Runs 2 and 3, the alloys were contained in uncovered high-purity iron crucibles. Alloys formed in the uncovered iron crucibles had to be chemically analyzed owing to the slight volatility of lead under the imposed experimental equilibration conditions. The compositions of these alloys were determined by a gravimetric analysis of lead as lead chromate. To avoid this problem, the remaining runs were made in covered graphite crucibles machined from high-density (1.92 g per cu cm) graphite rods (Basic Carbon Corp., Sanborn, N.Y.) which had a maximum ash content of less than 0.04 pct. The non- i reactivity of graphite with lead and magnesium has been previously established,' and the experimental procedure has been described in full detail.' In the runs using covered graphite crucibles it was found that the weight increases of the crucibles and their contents after equilibration equaled the magnesium contents of the alloys within the accuracy of the chemical analysis (i0.3 at. pct Pb). Clearly, the small lid openings of the graphite crucibles prevented significant lead losses during a run, and the weight increase was therefore taken as the magnesium content of the alloys. EXPERIMENTAL RESULTS At thermodynamic equilibrium the magnesium partial pressure of the specimens equaled that of the reservoir. Since the magnesium reservoir was always maintained at the temperature minimum of the system, the activity of magnesium of any specimen was given by the ratio of the magnesium reservoir pressure to the vapor pressure of pure magnesium at the specimen temperature. The standard state was chosen as liquid magnesium and the vapor pressure of liquid magnesium

Jan 1, 1968

By D. O. LIVINGTON

(Presented by invitation at a meeting of the Spokane Local Section of the institute, Feb. 17, 1912.) THE Pilares de Nacozari mine is located in Sonora, 75 miles south of Douglas, Ariz. The town of Douglas is on the International Boundary and is the place at which the ores from the Bisbee mines are smelted. The Moctezuma Copper Co. owns the mines at Nacozari, and the copper-concentrates shipped to the Copper Queen smelter, at Douglas, make a good smelting-mixture with the Bisbee ores. The Moctezuma Copper Co. and the Copper Queen smelter at Douglas, as well as the railroad from Douglas to Nacozari, are owned by the Phelps Dodge Co. The ore-deposit occurs in the form of a large ellipse with a major axis of approximately 2,000 ft. and a minor axis of about 600 ft., the major axis bearing about 90 W. of N. The whole of this ellipse is more or less mineralized; the surface being principally an iron gossan with some occasional copper-stains. Below the oxidized zone, which is not more than 50 ft. deep, the minerals are pyrite and chalcopyrite, with occasional seams of chalcocite, the latter being rare. There appears to be no well-marked zone of secondary enrichment, the oxidized gossan over the greater part of the deposit changing suddenly to what is apparently the original unaltered ore. The copper-values are concentrated around the perimeter of the ellipse, and it is principally around this perimeter that the mining is done. The ore mined averages a little more than 3 per cent. of copper, with a small amount of silver, less than an ounce per ton. Some ore of considerably higher grade than this has been shipped, however, but the above average is of the mill-run. The ore is wider near the two ends of the ellipse than along the sides, and is mined in some cases to a width exceeding 200 ft. The country-rock for the first vertical 500 ft is a volcanic acid breccia, probably rhyolitic; below

Sep 1, 1912

By N. S. Bornstein, M. A. DeCrescente

ThE nickel and cobalt base superalloys employed in gas turbine engines under some conditions' oxidize at an accelerated rate when exposed at elevated temperatures to atmospheres containing minor quantities of sea salt. It is generally agreed by investigators in the field that the major steps comprising this phenomena of accelerated oxidation, also called hot corrosion or more commonly, sulfidation, are a) a slag composed primarily of Na2S04 is formed on turbine components, b) the protective oxide scale is rendered ineffectual, c) the slag reacts with. the substrate to form metal sulfides, and d) the substrate is rapidly oxidized. Sodium sulfate, a minor constituent of sea salt, is formed within the gas turbine engine from the reactions between NaC1, the major constituent of sea salt, and the oxides of sulfur.' Depending upon concentration, engine compression ratio, and temperature, the Na2S04 will be present in the gaseous or condensed state. It has been determined that in order for sulfi-dation to occur the Na2S04 must be present in the condensed state.' Although a number of sulfidation mechanisms have been postulated, there is no consistent explanation of oxide scale breakdown and subsequent catastrophic oxidation. Some investigators believe that after oxide scale breakdown, the Na2SO4 reacts with the substrate to form lower-valent sulfur compounds. The accelerated rate of oxidation is due to oxidation of a nickel-sulfide phase." Other investigators believe that sulfur from the fused salt reacts with chromium in the alloy to form chromium-rich sulfides; thereby depleting the alloy of an element which imparts oxidation resistance.3,4 This publication will demonstrate that the accelerated rate of oxidation associated with sul-fidation is due to interactions between Na20, a con-stituent of NazSO.,, ancl the substrate. The aim of this study was to separate and determine the effect of compounds of sulfur and sodium on the oxidation kinetics of various nickel-base superalloys, in order to elucidate the sulfidation mechanism. The work reported herein is sponsored in part by the Naval Ship and Development Center-Investigation of EXPERIMENTAL PROCEDURE The nickel base alloys B-1900, U-700, and Waspaloy were chosen for this study. The nominal compositions are presented in Table I. The specimens, approximately 1 in. by 1 in. by 0.060 in. received in the "as cast" condition were polished with 600 grit emery, washed, and rinsed with acetone immediately prior to oxidation experiments. Sulfida-tion attack was duplicated in the laboratory by apply-ing a saturated aqueous solution of Na2S04 onto the al loy substrate and subsequently exposing the specimens in oxygen at elevated temperature. The method of Na2S04 application is described elsewhere in detail.6 In another set of experiments the affected substrate microstructure associated with sulfidation was reproduced free of the normal oxide scale, by reacting and diffusing elemental sulfur into the alloy. The experimental procedure is also described elsewhere in detail. All oxidation studies were performed using an Ainsworth thermobalance type RV-AU-1, which is readable to 0.01 mg and is reproducible to ±0.03 mg. The specimens, introduced into a heated furnace, were exposed in the temperature range 800" to 1000°C

Jan 1, 1970

By John C. McDonald

This report is a continuation of an earlier one with a similar titlell to which the reader is referred for such details of procedure as do not appear here. A brief summary will be given of the objects and methods of the research at this time. It was pointed out in the previous paper that properties of magnesium alloys, as reported by different workers, often differ. This was thought to be due to differences in the preparation of the specimens tested. The ideal way to compare alloys would be to prepare each of them in so many ways that reasonable assurance would exist that the best properties of each alloy had been developed. The alloys would then be compared on the basis of these properties. The treatment actually given the alloys of these studies is necessarily a compromise with completeness. Nevertheless, enough treatments have been given so that a considerable degree of assurance exists that the properties obtained are representative of the alloy as such, and not merely of the manner in which it was worked and heat-treated. Procedure The alloys were cast in the form of cylindrical ingots, which were extruded into bar before being rolled. Some of the ingots were made in 2-in. diameter molds whereas all of the alloys of the previous paper were cast in 4.in. molds. For the larger ingots, the extruded bar was 2 by 0.140 in.; lor the 2-in. ingots, it was I by 0.100 in. Preliminary work indicated that strip rolled from both sizes of the bars would have substantially the same properties. The bar was rolled at some temperature in the range 400' to 800°F., for a total reduction of 50 per cent. The temperature and reduction per pass were so selected that no cold cracking occurred, but were such as to leave the metal in a semihard condition. The hot-rolled metal was annealed at several temperatures in the range 400' to 800°F.; it was also cold-rolled, and given solution and aging treatments. Two standard test bars corresponding to each treatment were pulled in tension, and the average was obtained. Twelve treatments, in all, were given each alloy. The properties of the annealed specimen having the highest percentage of elongation were selected to represent the alloy in the annealed condition. The properties of the cold-rolled specimen having the highest yield strength were selected to represent the alloy in the cold-rolled condition. These properties were then plotted against percentage composition by analysis. Smooth curves were drawn among the points, thus graphically averaging the results over the entire range studied. The percentage of elongation in the cold-rolled state was not plotted, since it was always low, and varied in an irregular manner with composition. Alloys Studied Magnesium-calcium Alloys.—Fig. I presents the results on the alloys of the rnag-

Jan 1, 1941

By John C. McDonald

This report is a continuation of an earlier one with a similar titlell to which the reader is referred for such details of procedure as do not appear here. A brief summary will be given of the objects and methods of the research at this time. It was pointed out in the previous paper that properties of magnesium alloys, as reported by different workers, often differ. This was thought to be due to differences in the preparation of the specimens tested. The ideal way to compare alloys would be to prepare each of them in so many ways that reasonable assurance would exist that the best properties of each alloy had been developed. The alloys would then be compared on the basis of these properties. The treatment actually given the alloys of these studies is necessarily a compromise with completeness. Nevertheless, enough treatments have been given so that a considerable degree of assurance exists that the properties obtained are representative of the alloy as such, and not merely of the manner in which it was worked and heat-treated. Procedure The alloys were cast in the form of cylindrical ingots, which were extruded into bar before being rolled. Some of the ingots were made in 2-in. diameter molds whereas all of the alloys of the previous paper were cast in 4.in. molds. For the larger ingots, the extruded bar was 2 by 0.140 in.; lor the 2-in. ingots, it was I by 0.100 in. Preliminary work indicated that strip rolled from both sizes of the bars would have substantially the same properties. The bar was rolled at some temperature in the range 400' to 800°F., for a total reduction of 50 per cent. The temperature and reduction per pass were so selected that no cold cracking occurred, but were such as to leave the metal in a semihard condition. The hot-rolled metal was annealed at several temperatures in the range 400' to 800°F.; it was also cold-rolled, and given solution and aging treatments. Two standard test bars corresponding to each treatment were pulled in tension, and the average was obtained. Twelve treatments, in all, were given each alloy. The properties of the annealed specimen having the highest percentage of elongation were selected to represent the alloy in the annealed condition. The properties of the cold-rolled specimen having the highest yield strength were selected to represent the alloy in the cold-rolled condition. These properties were then plotted against percentage composition by analysis. Smooth curves were drawn among the points, thus graphically averaging the results over the entire range studied. The percentage of elongation in the cold-rolled state was not plotted, since it was always low, and varied in an irregular manner with composition. Alloys Studied Magnesium-calcium Alloys.—Fig. I presents the results on the alloys of the rnag-

Jan 1, 1941

By C. W. Robinson

Transportation is the minerals business. Once upon a time the geologist, the engineer and later the metallurgist reigned supreme, but the leading role in mineral development today is the economist-especially the transportation economist. Obviously this trend is more important for large volume, low value mine-to-market movements of such commodities as iron ore, phosphate, potash, coal and petroleum. But, transportation is of sufficient importance to all mining activities to be included in forward planning. Civilization has been influenced by transportation developments to an extent little understood or appreciated. Although industrialists are capable of intelligent historical analysis of this phenomenon, too frequently they have failed to project, or anticipate those changes in transportation techniques which will influence the pattern of industrial growth.

Jan 1, 1971

By O. D. Gonzalez

ThIS communication presents the results of a determination of the permeabilities of hydrogen and deuterium in a iron from 360° to 560°C. Recently Heu-mann and primas' have given values of the diffusivity and solubility in a iron of the two isotopes from which the permeability can be calculated. The work reported in the present communication, however, is a direct measurement of the rate of permeation of the isotopes through an identical iron specimen under the same conditions, so that systematic errors can be minimized in the calculation of the ratio of the permeabilities—the value which is usually of main importance in establishing limits for isotopic differences in diffusive processes. Experimentally the procedure used was the usual one: a steady-state flow of gas produced by a pressure difference across an iron membrane was observed as the pressure rise with time in a previously evacuated, calibrated volume connected to the outlet side of the membrane, while a constant pressure of gas was maintained on the opposite inlet side. The membrane was a cylindrical tube 9.38 cm long, 0.730 cm inside diam, 0.187 cm wall thickness, fabricated of 99.95 pct vacuum-melted National Research Corp. iron (less than 0.003 pct C). One end was sealed by a heliarc-welded iron (99.95 pctj lid of 0.147 cm thickness; the other end of the tube was welded to a 304 stainless-steel tube (0.318 cm ID and 0.318 cm wall) which connects to a vacuum system and to the calibrated volume including a dibutylphthalate manometer. The temperature of the iron membrane was maintained constant to ±0.l°C and was measured with an iron-constantan thermocouple which was calibrated against a standard Pt/Pt-Rh thermocouple: the temperature variation along the length of the membrane did not exceed ±0.5°C. Hydrogen was allowed to pass through the membrane assembly for 72 hr at 615°C before permeability measurements were begun. The system was tested at temperature before and after the experiments by a substitution of helium for hydrogen in the inlet side; under these conditions the pressure rise observed in the outlet side was negligible at all times. The gases used for the permeability measurements were 99.95 pct Matheson prepurified H2 and 99.7 pct D2 (0.3 pct HD) from the Isomet Corp. The gases were dried with liquid N2 before use. In Fig. 1 the logarithms of observed flows, jt, for a pressure difference of 73.4 cm Hg are plotted against T-1 Each point in Fig. 1 represents the slope determined by a least-square evaluation of a plot of the pressure in the outlet side of the membrane against time, including at least two separate sets of observations. Each point in the figure has an uncertainty of ±0.6 pct arising from the experimental uncertainty in the pressure and time measurements. The standard deviation of the observed pressures from the straight lines in all cases is less than the experimental precision of the measurements. The observed flow, jt, includes a small but not negligible flow through the stainless-steel connection.

Jan 1, 1968

By Winston W. Spencer

ALTHOUGH by far the largest A consumer of platinum metals in the world, the United States until recently has been in- significant as a producer. Writing in the "Minerals Yearbook" for 1939, H. W. Davis states that 1938 probably marked the beginning of an- other period of transition in the world production of the platinum metals. In that year, he says, their production in the United States jumped to 49,380 oz., or fourth in the rank of world output. In the ten-year period 1925 to 1934 United States output ,averaged only 8300 oz. Practically all of this sudden in- crease in production is attributable to the introduction of mechanized equipment at the placer deposits in the Goodnews Bay district of Alaska. This district adjoins the Bering Sea and Kuskokwim Bay (accented on the first syllable), and is approximately 120 miles south of the mouth of the Kuskokwim River. The nearest village and post office is Platinum, which lies about eleven miles north- east of the Goodnews Bay Mining Company's camp. A good gravel road connects the mine camp and the village.

Jan 1, 1940

By A. J. de Witte

INTRODUCTION The importance of determining dips of subsurface formations in wells, especially in wildcat wells, is self-evident and need not be enlarged upon. Various dipmeter instruments have been described in technical literature1,2,3,4 over the years. The introduction of the continuous dip-meters3,4 has made the earlier types obsolete. Continuous dipmeter logs are run by at least two different major service companies. Included as part of the service are a limited number of dip calculations—usually 10—made at different levels from the log. Very frequently this is the total result of a dip survey. It is believed that dipmeter surveys would be of considerably more advantage if geologists and engineers were able to make their own calculations from the original records. Determination of dips from the dipmeter record starts with the correlation of three log curves taken at different sides of the borehole. It cannot be over-emphasized that the quality of the results hinges primarily on this correlation. Nobody should be better qualified to correlate logs than a subsurface geologist or field engineer. The next requirement is to have an easy yet accurate way of computing the dip from the correlated points. Service companies use a mechanical device for this purpose. Directions for a graphical method were given in a recent article by B. 0. Prescotts. This paper describes a different graphical technique which, in our experience, is simpler, faster and at least as accurate. The entire manipulation is carried out on the stereonet in a number of easy moves. The paper is divided into two parts. This first part contains the introductory sections and full explanation of the geometry involved. The second part reviews the reading of data from an actual field log and summarizes the practical rules for computation of dips, the whole process being illustrated by an example. Following the outlined procedure, one could calculate a good many levels in a relatively short time, thus being able in particular to check suspected anomalies such as fault zones or unconformities in as close detail as the log will permit. GENERAL EXPOSITION OF THE METHOD Prnciple of the Dipmeter The position of a plane in space is defined by the location of three points, not on a straight line. Referring to Fig. I, if a borehole traverses a bedding plane, we can determine the strike and dip of the bed by finding three points which are not in line along this plane. Such points as A, B, and C are obtained with the dipmeter. The device consists of three electrodes (or caliper arms)1, 2, and 3 spaced at 120 angles, as schematically indicated, whose readings are separately recorded (Curves 1, 11, 111, to the right). When electrode 1 passes the boundary of the bed, Curve I will kick, etc. From these curves, we know the relative distances of points A, B, C, measured along the axis of the hole. If A is at zero, then B and C are at levels d1 and d2 up the hole,

Jan 1, 1957