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By H. W. Mossman

Three aspects of magnetite smelting are discussed. The first is the working out of equilibrium conditions for eliminating sulfur. The second is the influence of magnetite solubility on the difficulty of tapping the reverb matte. The third is an approximation of the equilibrium conditions in the reverb gases which govern whether magnetite is mode or reduced in the reverb slag by these gases and by any iron sulfide in the slag. MAGNETITE has had a varied history in the Hurley Smelter since its start in 1939. Magnetite determinations on the smelter products are made regularly only on the monthly composite samples. Variations on the monthly averages are shown in Table I. Magnetite which drops from the slag and matte in the reverb has some slight bottom buildup which comes and goes, but no substantial accumulation from this source has been found at the end of a normal nine months' furnace campaign. However, there has been some low grade magnetite bearing material mixed with considerable A1,0,, which has slid down from the bottom of the sloping flue between the reverberatory furnace and the waste heat boilers. This accretion has required drilling and blasting near the skimming end of the furnace. The magnetite has interfered with tapping at times. When the smelter was first started, tapping trouble from magnetite was extremely severe. Increasing the reverberatory furnace temperature by putting in an air preheater and a Dutch oven has helped greatly, although there still is occasional tapping trouble. When the present series of physical chemistry articles on copper smelting started coming out in 1950, they were read with interest, but no immediate application was seen for them. Results of some laboratory work in 1952 aroused a much stronger interest in this physical chemistry. A series of melts was made on some converter slags, which had magnetite in very large grain sizes, with the object of reducing the grain sizes in the slag, as it was known that it was easier to handle in the reverb in that condition. Anything done in the tests greatly reduced the grain sizes—even in the controls, where nothing was done except melt the slag and cast it. There was more magnetite in the slags after the tests than before, and with wide variations. There were no obvious reasons lor much of what happened in these tests. Much of the base material published in English in this field was made available for study. Recalculations were made on many of the type problems, and part of the data was reduced to local temperatures and compositions. Explanations were found for what happened in the 1952 series of tests on converter slags, and the same principles turned out to be a description of much of what magnetite does in the reverb. This article is to present the results of that study, from the viewpoint of applying the technical material in definite numerical form to the operating conditions in both the converters and the reverberatory furnaces at the Hurley smelter. Table I. Magnetite Variations on Monthly Averages, 1939 to 1955 Pet Magnetite Lowest Highest Average Converter slag 13.6 43.3 25.4 Roverb slaa 2.7 20.9 8.7 Matte 28 15.9 98 In general it was found that magnetite is made or reduced in both the converters and the rever-beratory furnace, depending on variations of temperature, matte composition, and reverb gas composition occurring in ordinary plant operation. Within reasonable limits, the field conditions for formation or reduction can be predicted, and probably can be set up and maintained. Converter conditions affecting magnetite formation can be put into numerical values better than for the reverb from purely technical calculations. The converter can be operated so as to keep the magnetite in the slag down to between 12 and 14 pct and still give satisfactory life for the converter brick. This depends upon having converter flux available which will make a slag with a good separation without raising the temperature too high. In the Hurley reverb and others with similar conditions, it is likely that a compromise of conditions will give a reasonably good control of combustion and still keep the magnetite from building up on the bottom. This discussion consists of three main parts. The first is the working out of the equilibrium conditions in the converter for determining in which direction the reaction 3 Fe,,O, (s) + FeS (1) F? 10 FeO (1) + SO, will go under actual converter operating conditions. The second deals with the influence of the solubility of magnetite in the slag and matte in the reverb on the difficulty of tapping matte. The third is an approximation of the equilibrium conditions in the

Jan 1, 1957

By E. C. W. Perryman

On electropolishing, high-purity Cu-Sn and Cu-Sn-P alloys with more than 5 to 9 pct Sn were found to contain many grain boundaries with a ridge-and-furrow profile. The effect was not eliminated by solution treatment and was present at the new boundaries in recrystallized material. It could be produced by diffusing tin into the 3 pct Sn alloy and it is concluded that the effect is due to enrichment of certain boundaries in tin, or possibly in an impurity present in the tin, in solid solution. IN the hot working of tin bronzes, especially those with moderately high tin contents, considerable difficulty is frequently encountered from intercrystal-line weakness, which results in cracking at the edges in sheet on strip rolling and in the unsupported regions of the rod in rod rolling.' These difficulties are not at all pronounced with up to 3 pct Sn, but at 5 pct Sn impose serious limitations and at higher tin contents, especially if the phosphorus content is also high, they may become very severe. In some work carried out at the laboratories of The British Non-Ferrous Metals Research Association it was found that a notable lack of ductility in tension at elevated temperatures, associated with intercrystal-line failure, was characteristic of tin bronzes even of relatively high purity. That the intercrystalline fracture is not due to a brittle or molten impurity phase at the grain boundaries was shown by the fact that the temperature at which intercrystalline fracture first occurred decreased with decrease of rate of straining. Tensile tests on some commercial tin bronzes, reported by Goetzel,² show a similar fall in ductility with increasing temperature. As no visual evidence had been obtained showing that the grain boundaries of the tin bronzes were different from those in copper, a more detailed microscopical examination was carried out, and the results of this are given in the present report. Several workers"' have observed that electrolytic polishing is a powerful tool for detecting small amounts of a second phase and in particular for detecting high solute concentrations, "equilibrium segregation," at grain boundaries. Since it had been found that the suggested solutions for electropolishing bronze and the standard solutions for brass and copper did not give satisfactory results it was first necessary to find a suitable electropolishing solution.' Experimental Procedure Materials Used: The cast and wrought materials used for this work were the same as those used for the mechanical tests referred to above. They comprised 3, 6, and 9 pct Sn alloys without phosphorus and the same three basic alloys with 0.03 and 0.4 pct P. A 5 pct Sn alloy without phosphorus, of similar purity, was also examined. The basis materials were Chempur tin, high purity 15 pct phosphor-copper and cathode copper, and spectrographic analyses are given in Table I. Treatment: The cathode copper was first melted under charcoal, degassed by plunging marble chips beneath the surface, and cast through a coal gas flame into sand molds to produce a number of small ingots which after machining were of a suitable size for subsequent vacuum melting. Chill cast bars of the alloys, approximately 1¼ in. in diameter, were then made by melting and casting in vacuo, the tern-

Jan 1, 1954

By A. D. Wilkinson

A CROSS-SECTION of the converter aisle through one of the converters and reverberatories is shown in Fig. 5. The aisle is 280 ft. long by 60, ft. wide by 681 ft. to the roof trusse, which have a 63-ft. span and are us )ported by steel columns spaced at 2b-ft. centers. Two 80-ton electric trails, supplied by the Welman Smith .Owen Engineering Co., with 20-ton auxiliary lifts, serve this building. Matte shells and converter-slag shells are dumped into heavily con¬structed flat-bottom bins located on the south side of the converter aisle and 1.0 ft. above the floor. The material is broken by ball and by hand, shoveled off into ladles for unloading on to the reverberatory charge floor, and charged into the furnace as previously mentioned.

Jan 1, 1934

By M. Earl Volin, M. Adnan Goksel

The hydro thermal agglomerating process described (U.S. Pat. No. 3,235,371, 1966)* is different from the cotzventional high-temperature processes. Green pellets or briquettes made with mixtures of moist iron-ore concen- trates and burned or hydrated lime fines are hardened in an azttoclave at 198°C and under steam pressure for 112 to 5 hr, with or without previous drying. Under the neutral conditiotzs, oxidation of magnetite does not occur as in the conventional processes. The resulting azzlomerates-- are hard and resistant to abrasion and weathering, and they do not disintegrate in the Under test. The equipment required for the process is not complicated, and favorable operating costs are projected because of the low temperamre and efficient heat recuperation. A low capital investment makes the process attractive even when the annual pellet production is below a million tons.

Jan 1, 1970

By Arthur Winslow

Location.—The Liberty Bell gold-mine is in the San Juan region, in the southwestern corner of Colorado. It is situated about 2 miles from the town of Telluride, near the summit of the Uncompahgre range, at the head of Cornett basin, at an altitude of 11,000 feet, and in a district of excellent record in the past and of great promise for the future. In the immediate vicinity are such well-known and important mines as the Smuggler Union, with a total production of over $12,000,000; the Virginius and Revenue Tunnel group, with a production of probably $5,000,000; the Tom-Boy, with a production of probably at least $3,000,000; the Japan, the Cimarron, and a number of others. History.—The first prospecting in this district is reported to

Jan 1, 1900

By Dale C. Matthews, Donald L. Simpson

Hecla Mining Company's Lakeshore Mine and Metallurgical complex near Casa Grande, Arizona, is the first plant in North America to utilize the roast/leach/electrowinning process for production of cathode copper from sulfide concentrates. Oxide copper ores are also mined at Lakeshore. Using sulfuric acid and sponge iron by-products of the roast/leach operation, the oxide ore is leached in vats, the copper is precipitated with iron and recovered as cement copper. This paper will describe the roast/leach facilities and discuss the operating conditions at which more than 95 percent of the copper in concentrate is made soluble in acid solutions while less than 5 percent of the iron is rendered soluble. More than 97 percent of the sulfur contained in concentrate is captured either as strong sulfuric acid or as soluble sulfates in leach liquors. The iron rich residue after leaching is pelletized and reduced to sponge iron in an SLRN reduction kiln.

Jan 1, 1981

By R. Crnojevich, J. L. Blanco, C. Nikolic, W. G. Sherwood

This paper describes the process used for cobalt purification and recovery at the AMAX Nickel Refining Co. plant In Port Nickel, Louisiana In the first stage, cobaltic precipitate produced in the Outokumpu electrolytic nickel cobalt separation process is dissolved using methanol and sulfuric acid liquor produced in this step is converted to an ammoniacal complex, and sufficient ammonium sulfate is added to facilitate cobalt oxidation to cobaltic pentammine Nickel is then selectively precipitated as nickel-ammonium sulfate double salt at pH 2 5 Final purification of the pentammine liquor is effected by pH adjustment with ammonia and removal of divalent nickel by ion exchange Purified liquor is then advanced to hydrogen reduction to produce cobalt metal Although similar to other cobalt refining processes used in the nickel industry, the Port Nickel process is unique because of a single point chemistry adjustment for a number of unit operations that occur in succession.

Jan 1, 1981

By J. E. Gruzleski, O. Knacke, M. J. Meixner

The evolution of CO from small specimens of solidifying steel was investigated using a thermobalance technique. Gas evolution begins in the temperature range between 1400" to .1450°C. This behavior is independent of the initial carbon content and the cooling rate. At a constant temperature above 1220°C, the rate of gas evolution is temperature independent. Out-gassing of the specimen is thought to proceed through small channels formed by the coalescence of bubbles between dendrites. LIQUID metals have a higher solubility for gas than the solid. During freezing, these gases form bubbles which may either escape or remain trapped in the solid. Recent work1 on the evolution of oxygen from solidifying silver showed that the gas is not liberated until a large fraction of the melt has solidified, and that the gas is evolved in an almost explosive manner. Because of the commercial importance of CO evolution in solidifying steel, a similar investigation was undertaken to study the mechanism of gas liberation in steel. EXPERIMENTAL The thermogravimetric method of Knacke and eixner1 to follow gas evolution by weight changes was employed. Weight differences of 0.1 mg could be reproducibly measured on specimens weighing 2.5 g. These specimens were taken from alloys with varying carbon contents from <0.01 to 0.11 wt pct and <0.01 wt pct of each of silicon, manganese, and aluminum. Specimens contained in 8 mm ID magnesium alumi-nate crucibles were heated to 1550°C under 1 atm of very pure (99.995 pct) CO with a tungsten resistance furnace. The specimen temperature was measured with a Pt-Pt/Rh thermocouple located directly beside the crucible. It is felt that the temperature thus measured was within 5" to 10°C of the specimen temperature since on fusion the weight of the specimen increased due to uptake of CO, and this weight increase was always observed to begin within the temperature range of 1530" to 1540°C. Even after 3 hr there was a small weight increase with time indicating that the molten specimen was not completely saturated with CO. Therefore, complete saturation was not attempted, and all specimens were simply held for 3 hr at 1550°C before cooling. Two types of cooling experiments were performed. One consisted of continuous cooling at linear rates from 5°C to 20°C per min down to 1020°C. The second type of experiment involved continuous cooling at 10°C per min to a predetermined temperature, and then holding the specimen at this temperature for a certain time. During all of these experiments, the sample weight was recorded as a function of time and temperature. OBSERVATIONS Continuous Cooling Experiments. No gas evolution was detected above 1450°C. The initial weight decrease always occurred between 1400" and 1450°C with the average at 1430°C. With further cooling, gas evolution continued to about 1150°C. Between this temperature and 1020°C no further outgassing was observed. This behavior did not seem to depend on the initial carbon content nor on the cooling rate in the range investigated. Typical patterns of gas evolution are shown in Figs. 1 and 2 for different initial carbon levels and cooling rates. The total amount of gas liberated during solidification varied with the rate of cooling. The total weight change on continuous cooling at a rate of 10°C per min

Jan 1, 1970

By S. Ramachandran, L. A. Vazquez, N. L. Grauerholz

A novel process is described for the selective flotation of scheelite from low-grade (- 0.5% W03) tungsten ores having appreciable quantities of calcareous and siliceous gangue minerals. Specifically developed for the Tempiute, Nevada facility of Union Carbide, the process yields tungsten concentrates exceeding 15 percent W03 with at least 85 percent recovery. This condition was not achieved through conventional processes. This process consists of conditioning the ore pulp with judicious amounts of lime, followed by soda ash and sodium silicate as modifiers. The scheelite is then floated using a mixture of oleic and naphthenic acids as collector. Contrary to the common observation as to the deleterious effects of calcium in fatty acid flotation systems, lime addition is found to be beneficial in achieving satisfactory results. The process has also shown excellent results on a Brazilian scheelite ore.

Jan 1, 1976

By G. Sachs, C. C. Chow, L. J. Klingler

Most structural parts which are heat treated are designed using strength properties which have been determined in the principal direction of the wrought material. For example, for rolled or drawn materials, properties are given for the rolling or drawing direction. The structures, however, may be loaded so that the critical stress is in some direction other than that for which the properties of the material are known. Investigations of forged products,123 have shown that while the yield strength and tensile strength of carbon steel billets and bars vary little with the direction of the test specimen in relation to the fiber, the contraction in area in tension tests and the impact strength in notched bar impact tests decrease in the transverse direction. The contraction in area and, consequently, the fracture stress of hard aluminum and magnesium alloy forg-ings have also been found to be lower in the transverse direction. An investigation on aluminum alloy plate4 likewise has shown the dependence of the fracturing characteristics upon the direction* of the test specimens, the longitudinal direction being considerably stronger than the transverse direction and the normal direction, with the normal direction being the least strong. The variation of properties with direction has been explained by a type of anisotropy called mechanical anisot-ropy. This anisotropy results from the elongation, in the direction of the principal strain, of certain phases, inclusions, and/or cavities in the metal during working. A mechanical fibering is thus produced which seems to persist through annealing and heat treatment. This investigation was initiated to determine the flow stress and fracture stress, at high hardness levels, at 90° to the rolling direction in a round steel bar. It is this direction which receives the critical stress in drawing dies machined from round bars. Preliminary tests showed a large difference in properties between the 0° and 90° directions. Consequently, it was felt that a more complete investigation, utilizing several types of tests, was warranted to determine the flow and fracture characteristics of a steel at various orientations, for a number of hardness levels. This investigation was conducted on an air hardening nondeform-ing die steel. Material and Procedure The distribution of properties was made on a 3-in. round bar of annealed high-carbon, high chromium steel of the following analysis: Pct Carbon................... 1.53 Manganese................ 0.39 Silicon.................... 0.27 Chromium................ 11.76 Vanadium................. 0.25 Molybdenum.............. 0.81 The 3-in. bar was produced from an 8-in. ingot, which was annealed and forged to a 4-in. square billet. The billet was annealed and rolled to a 3-in. round which was then annealed and straightened. This steel is an air hardening die steel which has very good dimensional stability on hardening; therefore, the residual stresses resulting from hardening would be expected to be low. A hardness survey across the diameter of the annealed bar showed no difference in hardness from the center to the outside. However, the test sections of all the specimens were taken approximately half way between the center of the bar and the surface to avoid any surface effect or possible porosity at the center. Tension, compression and bend tests were made on specimens hardened and tempered at six different temperatures. The tension test specimens, Fig 1, were machined from the bar at orientations of 0, 22.5, 45, 67.5 and 90° from the axis of the steel bar. The specimens were rough machined, heat treated and then ground to size. The test section on each specimen was lapped in a direction parallel to the axis of the specimen to remove any transverse scratches which might act as stress raisers. The specimens were tested in fixtures which insured concentricity of loading of less than 0.001 in.5 The transverse strains were measured with a radial strain gauge,5 the least count of which was 0.0001 in. change in diam. The compression specimens, Fig 1, were machined from the bar at ori-

Jan 1, 1950

By André G. Journel

Geological study of the Prony Nilaterite (New Caledonia) began in 1969 and included drilling, geological mapping and laboratory studies. The deposit is the product of weathering of basic rocks. From top to bottom, the following zones can be distinguished: a compact ferruginous cover; ore associated with yellow laterite, containing 1.0-1.5% Ni and low in impurities (SiO2, Mg); an intermediate silicated ore containing 1.3-2.0% Ni and higher percentages of impurities; garnierite ore much richer in nickel (2-4%) but high in SiO, and Mg contents (Mg up to 20%) ; and parent rock, largely peridotites. Distinction between types of ore is essential because the high grade ore (silicated and granierites) is penalized for its high Mg contents that demand additional acid treatment. All studies made distinguish two types of ore, laterite and silicated.

Jan 10, 1973

By H. T. Mulryan

For the producer of industrial rocks and minerals, marketing is often the only means of survival. His material must be sold. There is little of the impersonal force of the market regulating supply and demand. No commodity exchanges facilitate his efforts to dispose of the material he produces, and the success or failure of his business depends primarily on prompt and correct response to market needs and requirements. This response is particularly vital where materials of low unit value must be shipped to distant consumers owing to a relative rarity of suitable deposits. The advantage of positive response to change can best be illustrated by hypothetical cases showing how problems can be turned into opportunity through marketing ingenuity. Three such cases dealing with materials of unit values in the range of $5 to $100 per ton are outlined below.

Jan 7, 1969

By Thomas H. Clagett

ThE Pocahontas coal field comprises the area in Tazewell County, Va. and Mcrcer and McDowell counties, W. Va., in which Nos. 3 and 4, Pocahontas seams of bituminous coal are mined. It is a mountainous region with ridges rising 300 to 1000 ft. (91 to 305 111.) above the very narrow valleys and, except for some poorly kept farms along the tops of the ridges and the area occupied by the mining and coking plants and villages, is wooded throughout. First shipments were made from Pocahontas in 1883. There are now about 100 mines, 95 per cent. of which are drift mines. At the end of 1920, approximately 298,000,000 gross tons of coal had been produced and 35,000 acres of coal mined. No. 3 seam is from 4 to 10 ft. (1.2 to 3 m.) thick, but the change is gradual and the thickness fairly uniform in any particular locality. It has a streak of bone, about 2 in. (5 cm.) thick, to which the coal adheres on both sides. In parts of the field this bone is replaced by hard slate, usually from 2 to 4 in. thick. The seam has a slate or fireclay bottom and a draw-slate top. No. 4 seam is from 65 to 80 ft. (19.8 to 24 m.) above No. 3 seam, and of minable section only in the western part of the field. It is from 3 to 8 ft. thick and has two streaks of bone, cach from 1 to 2 in. thick. It has a fireclay or slate bottom and is overlaid by from 4 to 24 in. of a laminated coal and slate, locally termed "black rash," which is separated from the coal seam by about 3 in. of slate. This "black rash" averages about 2.5 per cent. ash, but contains some clean coal. The coal in both seams is quite soft. Cleavage sufficieiltly well defined to exert an influence on the mining is rare. Overlying strata arc slates, shales, and sandstones, the sandstones predominating, extending from 30 to 1000 ft. above No. 3 seam. The strata directly above the draw slate of No. 3 seam, or "black rash" of No. 4 seam, and up to the first sandstone are usually slates or shales, about 10 ft. thick over the.

Jan 1, 1923

By Thomas H. Clagett

ThE Pocahontas coal field comprises the area in Tazewell County, Va. and Mcrcer and McDowell counties, W. Va., in which Nos. 3 and 4, Pocahontas seams of bituminous coal are mined. It is a mountainous region with ridges rising 300 to 1000 ft. (91 to 305 111.) above the very narrow valleys and, except for some poorly kept farms along the tops of the ridges and the area occupied by the mining and coking plants and villages, is wooded throughout. First shipments were made from Pocahontas in 1883. There are now about 100 mines, 95 per cent. of which are drift mines. At the end of 1920, approximately 298,000,000 gross tons of coal had been produced and 35,000 acres of coal mined. No. 3 seam is from 4 to 10 ft. (1.2 to 3 m.) thick, but the change is gradual and the thickness fairly uniform in any particular locality. It has a streak of bone, about 2 in. (5 cm.) thick, to which the coal adheres on both sides. In parts of the field this bone is replaced by hard slate, usually from 2 to 4 in. thick. The seam has a slate or fireclay bottom and a draw-slate top. No. 4 seam is from 65 to 80 ft. (19.8 to 24 m.) above No. 3 seam, and of minable section only in the western part of the field. It is from 3 to 8 ft. thick and has two streaks of bone, cach from 1 to 2 in. thick. It has a fireclay or slate bottom and is overlaid by from 4 to 24 in. of a laminated coal and slate, locally termed "black rash," which is separated from the coal seam by about 3 in. of slate. This "black rash" averages about 2.5 per cent. ash, but contains some clean coal. The coal in both seams is quite soft. Cleavage sufficieiltly well defined to exert an influence on the mining is rare. Overlying strata arc slates, shales, and sandstones, the sandstones predominating, extending from 30 to 1000 ft. above No. 3 seam. The strata directly above the draw slate of No. 3 seam, or "black rash" of No. 4 seam, and up to the first sandstone are usually slates or shales, about 10 ft. thick over the.

Jan 1, 1923

The following list contains the names of members whose death notices have been printed in MINING AND MDTALLURY from April 5. 1944 through March 15. 1945, Biographical sketches published in MINING AND METALLURGY are indicated in the last two columns. [ ]

Jan 1, 1946

By F. H. Spedding

CONSIDERABLE confusion exists in the literature concerning the phase transformations exhibited by the lighter rare earth metals; namely, lanthanum,1-8 cerium,1,3,5,8,9-15 praseodymium,8 and neodymium.1,8,16,10 To obtain further information

Jan 1, 1958

By J. W. Putman, N. J. Grant, D. S. Bloom

SEVERAL recent determinations of the melting S point of pure chromium have been reported which give values of 1845°C1; 1895°C,² 1930°C,³ 1860°C,&apos; and 1890°C.5 because of this wide spread of values, it appeared desirable to make one additional attempt to obtain a more accurate, reliable figure. bloom and Grantx recently reported a phase transformation in chromium:.--- it was desired to check this, too, in the course of checking the melting point. To accomplish this, a much higher purity chromium was required. Electrolytic chromium containing about 0.5 pct 0 was crushed to a finer size and was then annealed in highly purified, dried hydrogen at 1375°C for 100 hr. This treatment produced chromium with about 0.008 pct 0, 0.002 pct N, and negligible carbon. The balance of the impurities was 0.3 pct Fe, 0.03 pct Si, 0.004 pct S, and less than 0.001 pct Mo. This chromium was melted in a stabilized zirconia crucible under pre-purified argon, using induction heating. Temperature measurements were made with annealed wolfram-molybdenum thermocouples and a Leeds and Northrup Speedomax Recording Potentiometer. The thermocouples were annealed for 1 min in hydrogen at a temperature of about 2400°C. Each leg of the thermocouple was immersed in an ice-water mixture using mechanical connectors to the lead wires. The thermocouple tip was immersed in the molten chromium in a zirconia protection tube which introduces an error of about 3°C as a temperature drop through the walls of the tube. The weight of chromium was about 183 g. The dimensions of the melt in the crucible were about 1.25 in. diam x 2 in. high. The tip of the thermocouple was held about ½ in. from the bottom of the crucible. Two new thermocouples were used with the above set-up giving the results shown in Table I. Since this chromium is purer than that used by Greenaway, Johnstone and McQuillan,&apos; by Carlile et al.,4 and other investigators,²,3 it is believed to be the more accurate. Thermocouple aging effects which are believed responsible for the high value of 1930°C³ were avoided in this work. This value of 1903°C is correct to about ±10°C. The right hand column lists the temperature of the a ? ß transformation on heating and the ß ? a on cooling.3 A sample curve is shown in Fig. 1. The value so determined is 1840°C + 15°C. Fig. 1 gives adequate evidence that there is a transformation in chromium at high temperatures.&apos; References &apos;H. T. Greenaway, S. T. M. Johnstone, and M. K. McQuillan: High Temperature Thermal Analysis USing the Tungsten-Molybdenum Thermocouple. Journal Inst. Metals (November 1951) 19, p. 109. ²J. W. Putman, R. D. Potter, and N. J. Grant: The Ternary System Chromium-Molybdenum-Iron. Trans. A.S.M. (1951) 43, p. 824. ³D. S. Bloom and N. J. Grant: Chromium-Nickel Phase Diagram. Trans. AIME (1951) 191, p. 1009; Journal of Metals (November 1951). &apos; S. J. Carlile, J. W. Christian, and W. Hume-Rothery: The Equilibrium Diagram of the System Chromium-Manganese. Journal Inst. Metals (1949) 76, p. 169. 5G. Grube and R. Knabe: Ztsch. Elektrochemie (1936) 42, p. 793.

Jan 1, 1953

By John H. Jett

MATERIAL and supply handling is a major problem at underground mines, especially to those operating through deep shafts with small compartments. Mechanization is necessary, yet in many cases cannot be utilized to fullest advantage owing to the limitations of existing plant design. In cases of this sort, improvisation is indicated and adds to the headaches and enjoyment which most mine operators get out of their work. Underground operations at the Copper Queen Branch are principally carried on through three vertical shafts: the Junction shaft, Campbell shaft and Cole shaft. Hoisting depths range from 1400 ft at the Cole through 2700 ft at the Junction to 2966 ft at the Campbell shaft. Men and supplies are moved through each of the shafts. Cage compartments are small requiring all material over 5 ft in length to be handled in an upright position on the cages.

Jan 3, 1953

By D. G. Peterson, W. J. Guay

The extraction of gold from refractory ores containing activated carbon and various types of carbon compounds is being done commercially for the first time at Carlin, Nev., by Carlin Gold Mining Co. Oxidation of the carbonaceous materials to CO and CO,- is accomplished by the introduction of chlorine into an aqueous pulp. The chlorine reacts with the limestone in the ore to form calcium hypochlorite which in turn reacts with the carbonaceous matter to form calcium chloride, CO, and CO.. Following the oxidation treatment the ore is cyanided in the conventional manner. For the first six months of operation the plant averaged 83% extraction at a feed rate of 500 tpd and a consumption of 28.4 lb chlorine per ton of ore which showed only 33.5% extraction by conventional cyanidation without the oxidation pretreatment.

Jan 1, 1974

By Samuel G. Lasky

The writer has been interested in determining the mineralogic dis-t,ribut.ion of silver in the base-metal ore of thc Ground Hog mine of the Asarco Mining Co. in the Central mining district of New Mexico.&apos; This ore consists of a varitextured mixture of sphalerite, chalcopyrite, galena and pyrite in a quartz gangue. The average ore contains 14 per cent of zinc, 9.5 per cent of lead, 5 per cent of copper, and 10 oz. of silver per ton. Generally it is accepted that the silver in ores of this type, in which no definite silver minerals can be recognized, is contained in the base-metal minerals. Published assays of pure mineral specimens from many districts show that each of the base-metal minerals contributes in some degree to the total silver content of the ores, and to determine the proportionate amount contributed by each mineral in any specific ore it has been the practice to assay selected specimens of each mineral, the purity of which has been determined microscopically. There being no such assays available for the Ground Hog ore, and it being impracticable to have enough of them made to give average results, the writer resorted to a mathematical attack upon composite assays of large quantities of ore and concentrates. Surprising results were obtained, and it is the threefold purpose of this paper to present these results, to describe the mathematical method used, and to point out the value of using this method even when assays of pure specimens are available. Value of the Mathematical Method The obvious way of attacking the problem would be to separate cleanly large quantities of ore into its component minerals and then to determine the average silver content of each mineral by assay of representative samples. Since an absolutely clean separation is not feasible, it has been the practice, as mentioned above, to assay pure specimens of the different minerals. No one will contend seriously that the silver content of a base-metal mineral is constant throughout any particular deposit, and it will

Jan 1, 1935