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By G. R. Love, M. L. Picklesimer

Aboue 950°C the Nb-Zr system consists of a completely miscible bcc solid solution, commonly called the phase. Between 950 and 600°C, and between 20 and 85 pct Nb, the phase deconlposes, after sunciently long times, into two bcc solid solutions. The pct Zr alloys are conveniently descibecl with T-T-T (time-temperature-transformation) curves having a nose at about 2 hr at 700°C. The reaction rate varies only slowly with zirconium content and negligibly with oxygen contanzination; it is speeded up by a factor of 10 to 15 by 90 pct cold ulork and slowed dou by n factor oj 10 to 30 by a two-hundrecljold increase in grain size. Nb-r alloys with compositions between 40 and 85 pct Nb have been the basis for the majority of commercially important superconducting materials. In part because of their commercial promise, more is known about these alloys than about most other high-field superconducting materials. At the same time, there is considerable disputed or incomplete metallurgical information. For example, although Rogers and tkins&apos; indicate a monotectoid reaction at approximately 600°C and a two-phase 01 + 0, field extending between 20 and 85 pct Nb and to a maximum of 95OGC, erhout&apos; has reported that this entire region would be a single homogeneous B were it not for oxygen contamination. Again, although it has been shown that relatively short-time heat treatments in the vicinity of 700CZ significantly improve the ability of short wire samples to carry high currents in high magnetic fields at 4.2K, these observations have never been fully correlated with the structural change or changes occurring during the anneal. We intend to investigate in detail the effect of metallurgical variables, including heat treatment, on the superconducting properties of hard superconductors. To verify that our experimental techniques are valid and to establish a relative standard against which other materials may be measured, we feel it advisable to know the behavior of the Nb-Zr alloys under a variety of processing conditions. As an initial step toward this goal, we have determined in detail the kinetics of the transformations in Nb-Zr alloys. EXPERIMENT A number of problems had to be solved before beginning any fruitful work on the reaction kinetics in this system. While solving some of these problems, either by chance or by design, small amounts of information were obtained about alloys containing 40, 50, 60, 65, 67, 70, and 75 pct Nb, bal. Zr. In addition, a large range of grain sizes and a range of temperatures considerably greater than the range indicated by Rogers and Atkins phase diagram were examined. We will, however, report in detail only the results obtained for the Nb + 33 pct Zr and Nb + 25 pct Zr alloys at three grain sizes, two levels of oxygen contamination, and the temperature range 550 to 950°C. These data are most complete, but the other data are sufficiently complete to indicate the kind and magnitude of the variation of the transformation kinetics outside this range. The first and most difficult problem encountered in this inquiry was one of sample homogeneity. When Nb-Zr alloys are arc- or electron-beam-melted on a cooled copper hearth, solidification is sufficiently slow that there is appreciable coring in the cast structure and a large variation of grain size across the button thickness. Both these factors significantly affect the apparent reaction rate in the system. A two-step solution to the problem was attempted; an arc-melting and drop-casting technique has been developed by conald that greatly reduces the as-cast grain size and virtually eliminates coring segregation. Ingots made in this way exhibited no detectable (3 pct maximum) zirconium segregation. Before it was evident just how good this technique was, we attempted to supplement it with rather long-time, high-temperature annealing of the cast ingots. This annealing was carried out in evacuated and sealed (seal-off pressures < 1.0 x 106 torr) quartz capsules lined with tantalum foil at 1400 to 1450 C for 8 to 72 hr. There were two principal effects of this treatment: the grain size increased to a fairly uniform 150 p, and the surface and all grain boundaries near the surface acquired a film of a second phase, tentatively identified as an oxide (possibly additionally contaminated with silicon). There was no evidence that this 1400 C treatment had affected the zirconium segregation. High-temperature annealing was subsequently used only for grain-size control, but anneals of longer than 4 hr at temperatures greater than 1000°C were performed in dynamic vacuums (pressure no greater than 1.0 x lo torr). Any contamination resulting from these treatments was well below the limits of detection of our techniques. All samples, as cast, were cold-swaged to at least 85 pct reduction in area. The samples called cold-worked were tested as swaged. The minimum re-crystallization anneal for these alloys was about 12 hr at 1050 C; this produced an equiaxed grain diameter of about 4 to 8 P. Annealing for 4 hr at 1450°C produced a grain size of about 80 to 150 p; and annealing for 4 hr at 1650aC, close to the melting point of many of these alloys, produced a grain size of 0.5 to 1.0 mm. At all temperatures, the larger grain size was

Jan 1, 1967

By W. C. Hellyer, R. A. Campbell

In single-stage autogenous grinding, the buildup of a critical size fraction in the media can be corrected by removing this material through pebble ports, crushing it below the critical size range, and recycling the crusher product back to the mill. The rate at which the critical sire fraction is crushed affects the size distribution of the grinding media and this in turn affects the sire distribution of the grate discharge. This provides a means for controlling the grind produced by a single-stage autogenous grinding unit. The pilot-plant investigations on a very hard copper ore were carried out at the facilities of the Institute of Mineral Research, Michigan Technological University. In an autogenous grinding circuit in which feed at approximately 9 in. top size is reduced to a size suitable for subsequent processing, the build-up of a "critical size" fraction in the media causes problems. A "critical size" fraction has been defined as, "media too small to effect reduction by impact grinding of ore coarser than a quarter of an inch and too large to be broken by the largest size of media in the charge."&apos; The buildup of a critical size fraction reduces the capacity of the mill, increases the grinding power requirements per ton of finished product, and generally produces a finer grind than is desired. It is general practice to overcome this problem by the addition of large-diameter steel balls to the grinding charge. This has certain disadvantages, such as 1) an increase in mill liner wear, 2) wear on the steel balls, and 3) some loss in flexibility in grinding circuit opera-tions resulting from difficulties in removing the steel balls by means other than by grinding out. A number of investigators&apos; have suggested the use of a small crusher as an alternative to the use of large steel balls for control of a critical size fraction. With this technique the critical size fraction is removed from the mill continuously through suitable-sized pebble ports, crushed below the critical size range, and returned to the mill. An external crusher would cause no increase in mill liner wear, the wear on the crusher would be less expensive than the wear on the steel balls, and the flexibility of the grinding circuit operations would be enhanced since the crusher can be cut into and out of the circuit at will. This paper describes an autogenous grinding pilot plant investigation on a very hard copper ore, which led to the selection of an autogenous grinding flowsheet incorporating a crusher in the circuit as the preferred method for grinding this ore. Further development of the technique demonstrated that the crusher could be employed to control the product produced by an autogenous grinding unit. Description of Pilot Plant These investigations were carried out at the pilot-plant facilities of the Institute of Mineral Research, Michigan Technological University, Houghton, Mich. The autogenous grinding unit was a 6-ft-diam by 2-ft-long Hardinge Cascade mill with %-in. slotted steel grates. Pebble ports cut into the grates allowed passage of pebbles having a top size of approximately 21/2 in. The grate and pebble-port discharge material passed over a 3/16-in, trommel screen with the trommel under -size being pumped to the classification device. DSM screens were employed for classification in the early investigations, but were replaced later by a small Dorr rake classifier. The trommel oversize material returned to the mill via scissor conveyors. When the crusher was incorporated into the circuit, the trommel oversize material passed to a double-deck vibrating screen fitted with 2-in. and 3/4-in. square mesh screen cloths. The $ 2-in. and the —2 +3/4-in. size fractions were combined and crushed to a nominal 1 in. in a 4 x 6-in. jaw crusher. The crusher discharge and the —3/4-in. screen undersize returned to the mill. In the final investigation, a flap gate fitted to the top deck of the screen directed the — 21/2 +2-in. size fraction to the crusher or back to the grinding mill. The flap gate operated manually on a 15-min cycle, with this material being crushed for so many minutes out of each cycle. This was found to be a more reliable method of controlling the amount of — 21/2 +2-in. material crushed than attempting to make a split of a small weight of material on a continuous basis. Operational Techniques: Each sample was sized into three fractions and the Cascade mill feed was reconstituted from these size fractions in the proportions existing in the original sample. Sufficient feed for 15 min operation was weighed out and fed by hand over a 15-min period. The gross mill power draft was recorded every 15 min, corrected for tare power and drive efficiency, and reported as net kilowatt-hours per ton. A recording kilo-watt-hour meter provided a continuous visual record of the power drawn by the mill. Pulp densities of the trommel undersize and the classifier overflow (or DSM undersize) were taken every 15 min. The classifier overflow was sampled automatically. Timed samples of the classifier sands were taken every 15 or 30 min, weighed, and returned to the circuit. After correcting for moisture content, the weights were converted into percent circulating load. Timed samples of the +2 in., the —2 + 3/4-in., —3/4-in. screen fractions, and the crusher discharge, were

Jan 1, 1971

By G. Derge, Ling Yang, G. M. Pound

The specific conductance and its temperature dependence were measured over the entire composition range of the molten Cu2S-CuCI system. At a typical temperature of 1200°C, 10 rnol pet of the ionically conducting CuCl reduced the specific conductance from about 77 ohm-lcm-l for pure Cu2S to about 32 ohm -1cm -1, and 50 mol pet CuCl reduced the conductance to that for pure CuCI—about 5 ohm 1cm1. The nature of electrical conduction in molten Cu2S, FeS, CuCI, and mixtures was studied by measuring the current efficiency of electrolysis at about 1100°C. The Cu2S, FeS, and mattes were found to conduct exclusively by electrons, but addition of 1 5 wt pet CUS to Cu2S produces a small amount of electrolysis. Addition of CuCl to Cu2S suppresses electronic conduction, and ionic conduction reaches almost 100 pet at a CuCl concentration of about 50 mol pet. These facts are interpreted in terms of electron energy level diagrams by analogy to the situation in solids. RESULTS of electrical conductivity studies on molten Cu-FeS mattes as a function of composition and temperature have been reported.&apos; The specific conductances ranged from about 100 ohm-&apos; cm-&apos; for pure Cu2S to 1500 ohm-&apos; cm-1 for pure FeS. This is in sharp contrast with the low specific conductance of molten ionic salts for which the transfer of electricity is by migration of ions in the field. In general, these ionically conducting molten salts, such as NaC1, KC1, CuC1, etc., have a specific conductance of the order of magnitude of 5 ohm-&apos; cm-&apos;. It was concluded on the basis of this evidence that molten FeS and Cu,S exhibit electronic conduction. Pure molten FeS has a small negative temperature coefficient of specific conductance, resembling metallic conduction, while pure molten Cu2S has a small positive temperature coefficient, resembling semi-conduction. The molten Cu2S-FeS mattes follow a roughly additive rule of mixtures, both with respect to specific conductance and temperature coefficient. Savelsberg2 has studied the electrolysis of molten Cu2S and Cu2S + FeS. He concluded that while molten Cu2S is an electronic conductor, there is some ionic conduction in molten Cu2S + FeS3 owing to the formation of the molecular compound 2Cu2S.FeS and its dissociation into Cu1 and FeS2-1 ions. The present work does not verify his results. Chipman, Inouye, and Tomlinson" have studied the specific conductance of molten FeO and report a high specific conductance, about 200 ohm-1 cm-1 of the same order of magnitude as that found for molten mattes, and a positive temperature coefficient. They interpret these results in terms of p-type semiconduction in the ionic liquid by analogy to the situation in solid FeO.1 imnad and Derne&apos; detected appreciable ionization in molten FeO by means of electrolytic cell efficiency measurements. In order to verify the conclusion that electrical conduction in molten Cu2S and mattes is electronic, and to gain further insight into the structure of molten sulfides, the following investigations were carried out in the present work: 1) The specific conductance, s of the molten system Cu2S-CuC1 was measured as a function of temperature over the entire composition range. As discussed later, molten CuCl is an ionic substance. It was thought that if molten Cu2S were simply ionic in nature, addition of small amounts of CuCl might not have a catastrophic effect in lowering the high conductance of the Cu2S. On the other hand, if much electronic conduction occurs, addition of the ionic CuCl should have a large effect in destroying the electronic conduction. 2) The electrolytic cell efficiency of the following molten systems was measured at about 1100°C in specially designed cells: Cu3; Cu2S + FeS, 50:50 by wt; FeS; Cu2S + CuS, 15 wt pet; Cu2S + CuC1, 5.9 to 46.4 mol pet; and CuC1. This gives a direct measure of the fraction of current carried by ions in these melts. Further, the cell efficiency, extrapolated to zero ionic current, is given by cell efficiency = (s leasile + s elexstronic). [1] s lucile for molten CulS would be expected to be no greater than that for molten CuC1, whose s lonle is about 5 ohm-&apos; cm-1, as will be seen in the following. u,.,,.,.,.......for molten Cu,S is of the order of 100 ohm-&apos; cm-&apos;.&apos; Thus, a large increase in cell efficiency from 0 to values of 10 to 100 pet upon addition of CuCl to Cu2S would indicate destruction of the electronic conductance. Conductance Measurements Experimental Procedure—The apparatus and experimental method were the same as those described in detail in connection with the study of electrical conduction in molten Cu,S-FeS mattes.&apos; A four terminal conductivity cell and an ac poten-

Jan 1, 1957

By Hsun Hu, B. B. Ruth

The rates of migration of high-angle boundaries in zone-refined aluminum crystals rolled 20 to 70 pct in the (110)[i12/ orientation were studied. Following a recovery anneal at an appropriate temperature to stabilize the polygonized structure, boundary migration rates of artificially nucleated grains were measwed isothermally at several temperatures. Results indicate that the rate of boundary migration depends strongly on the amount of deformation and on the cell size of the polygonized matrix, and is related to the driving free energy by a power function. The degree of anisotropy in growth 0.f the re crystallized grains nn&apos;th preferred mientation is independent of deformation; the migration rates of the fast-moving and the slow-moping boundary segments of a gowing grain differ by as much as one order of magnitude. The actir\ation energy fm a grain boundary migration, although nearly the same for both the fast-moving and the slow-moving boundaries for a given deformalion, decreases from 45 to 30 kcal per mole with an increase in deformation from 20 to 70 pct reduction. Re crstallization by the growth of the artificially nucleated grains results in preferred orientation. The Percentuge of&apos; grains favorably oriented for growth increases with increasing deformation. None of these grains corresponds to the ideal Kronberg-Wilson orientation relationship. The observed growth aniso-tropy is discussed in terms of boundary structure. The boundary velocity as a function of the cell inter -facial area, or the driving free energy, is discussed in the light of current theories of boundary migration. It is well established that recrystallization with re-orientation occurs by the migration of high-angle boundaries of strain-free grains. The driving force for this process is provided by the free energy stored in the metal during deformation. A quantitative study of the effect of varying driving force on grain boundary migration in deformed metals has not been possible heretofore, primarily because of: 1) concurrent recovery steadily decreasing the available driving free energy for boundary migration, &apos;-3 and 2) in-homogeneity of strain in the deformed metal.4 Aust and Rutter3 studied grain boundary migration in striated single crystals of zone-refined lead. Although the driving free energy in such crystals remains unaltered during annealing, this method does not provide a range of driving free energies over which measurements of grain boundary migration can be made. In the present investigation, the rates of migration of high-angle boundaries in deformed aluminum zone- refined single crystals were studied at various temperatures, after deformation ranging from 20 to 70 pct reduction by rolling at -78°C in the (ll0)[i12] orientation. The boundary migration rates along different crystallographic directions were determined under steady-state conditions, i.e., in the absence of competing recovery processes or impingement of recrystallized grains growing into the deformed single crystal matrix. Simultaneous recovery was eliminated by suitable anneals prior to the boundary migration measurements. The recrystallized grains, which grew a ni so tropically into the homogeneously polygonized matrix, developed flat boundary segments during early stages of growth. These boundary segments subsequently migrated along a direction approximately normal to the boundary plane into the matrix rystal. Increasing deformation over the range employed was estimated to increase the driving free energy for boundary migration by about five times. The kinetics of the boundary migration process, examined under these conditions, indicate that the boundary velocity is greatly affected by a small change of the driving free energy in the matrix crystals. These results were examined in the light of the current theories of grain boundary migration. EXPERIMENTAL PROCEDURES Single crystal strips (9 by 1 by 0.125 in.) of zone-refined aluminum, were seed-grown by the Bridgman method in a high-purity graphite mold (<lo ppm ash) at 1 in. per hr. Precautions were taken to minimize contamination of the metal during crystal preparation and subsequent handling. Spectrographic analysis of the metallic impurities in the grown crystals is Qven in Table I. The crystals were rolled in the (110)[112] orientation at -78°C to various reductions in thickness, ranging from 20 to 70 pct, in 10 pct increments. The desired reduction was achieved by many rolling passes, each being no more than 0.002 in. To minimize surface friction, the crystal was rolled between two thin layers of teflon. For those crystals rolled more than 40 pct, it was necessary to remove the disturbed surface layers by electropolishing at -5" to -10°C at an intermediate stage of rolling. The edges of deformed crystals were removed by a jeweler&apos;s saw while submerged in alcohol at -78° C to obtain samples of about ? by i in. The distorted metal at the cut edges and the surface layers were then removed by electropolishing, with removal of a minimum of 0.004 in. from each surface. The thickness of the crystals prior to rolling was chosen so that the final thickness was 0.025 in. for all samples. These deformed single crystals were each prean-nealed for 1 hr at an appropriate temperature in the range of 130" to 280°C, depending upon the amount of deformation. The purpose of this preannealing was to

Jan 1, 1970

By F. V. Greene

This paper is based on my experience in the use of asphalt, for paving and other purposes, during the last ten years, part of the

Jan 1, 1889

By A. Eilers

I PRESENT herewith, for the inspection of the members of the Institute, a specimen of coke, made in gas-retorts from the lignite of Trinidad, Colorado.

Jan 1, 1874

By A. Eilers

I present herewith, for the inspection of the members of the Institute, a specimen of coke, made in gas-retorts from the lignite of Trinidad, Colorado.

By Clarence R. King

Diamond drilling of Cove Creek deposits and development of flotation-refining process may lead to commercial operation of property worked sporadically for almost a hundred years.

Jan 4, 1953

By T. R. A. Davey

A Survey of current and recent research and developmental work is supplemented by predictions for processes which should be developed over the remainder of this century.

Jan 1, 1970

The authors have done a very thorough job in analyzing the factors affecting turndown sulfur performance at Inland&apos;s No. 4 B.O.F. While many of the results are not unexpected, it is helpful I am sure to have quantified many of the relationships that the operators rely upon daily. On the other hand, however, the results shown for ladle skimming are very surprising. I do not doubt the statistical relationship shown but I feel, based on our experience at Youngstown, that some other parameter has been ignored in the data on ladle skimming. It was our experience that by skimming our hot metal transfer ladle we could reduce our lime consumption by eleven percent while obtaining sulfur removal comparable to that shown by the authors. I would like the authors comment of similar adjustments on the lime charge that were made on skimmed heats at Inland. If not, it would seem that one could gain more desulfurization in the B.O.F. by raising the hot metal manganese than by installing a skimmer. The author&apos;s state that no comparisons are valid between the data sets for the two time periods involved because of the difference in the consumption of No. 2 Bundles. I feel this in fact is one of the important conclusions to be drawn from the paper, as there seems to be approximately a 10-20 percent difference in percent of heats that are under 0.020 percent sulfur for similar conditions, which is an extremely important difference with today&apos;s demands for low sulfur. I would like the author&apos;s comments on the difference in the amount of 82 bundles consumed. Along this same line I was quite surprised by the relatively low number of heats that had a first turn down of 0.020 percent sulfur or less. Also rather disconcerting are the conditions necessary to achieve a first turn of 0.020 percent sulfur, 50 - 60 percent of the time, as required by today&apos;s specifications. It is a problem which we are well aware of, but the numbers are still shocking.

Jan 1, 1972

By William M. Newton, John E. Sherborne

The value of the electric log as a means of interpreting underground structures has been increasingly demonstrated by its almost universal present-day use. It becomes important, therefore, to be able to evaluate the factors that affect the values obtained. By far the greatest number of electric logs are measured today in boreholes filled with rotary mud, the properties of which may exert an important influence on the electric log. In an effort to establish what the order of magnitude of such effects might be, a series of investigations was made in the laboratory and the field. Apparatus suitable for either field or laboratory use was designed and a procedure established for determining the factors that affect the resistivity of drilling muds over a wide variation in mud condition. Five muds representative of the types of clay-base muds employed in California were tested and the following observations were made: 1. The effect of raising the temperature from 8o° to i8o°F. is to decrease the resistivit~ of the mud or filtrate approximately 50 per cent. 2. The resistivity of the mud in most cases closely approximates that of its filtrate. 3. The change in the resistivities of muds caused by the addition of chemical is not the same function of the amount of chemical added for each mud. 4. The effect produced by increased sodium chloride content is to reduce markedly the resistivity of the mud and its filtrate. 5. Weighting materials, such as baroid and limestone, tend to increase the resistivity of drilling mud. 6. Cement and counteracting reagents reduce mud resistivity. The results of the investigation indicate that the mud resistivity is materially affected by the addition of electrolytes whether these are added for the purpose of conditioning the mud or enter the mud when salt beds or highly saline structures are penetrated. It should be observed, however, that the magnitude of the effect produced upon the electric log by changes in mud resistivity is also dependent upon the relative salinity of the water en. countered in the formation being logged. The laboratory data are supplemented by field data, which indicate that both the self-potential and resistivity curves are influenced by mud resistivity.

Jan 1, 1946

By Roy A. Hardy

THE Getchell mine is a comparatively recent discovery in the old Potosi mining district, Humboldt County, Nevada, a district organized in the seventies and eighties, in which some prospecting was done in several metamorphic zones along a monzonite-sedimentary contact. The district had no production in the early days and no further work was done until the same metamorphic zones were prospected for tungsten (scheelite) in 1917 and 1918, when a few small lots were shipped. It again lay dormant until 1934, when prospectors investigated a large siliceous escarpment at one end of the old district where outcroppings were a prominent feature of the topography. Undoubtedly, during the early prospecting of the district, the outcroppings had been investigated for mineral, probably by panning, with negligible results because the gold later found is invisible in the concentrate and is refractory. A fire assay revealed the presence of sufficient gold to warrant development work. The outcroppings were easily prospected by short tunnels having their portals in the siliceous material on the hanging-wall side, with a maximum distance to crosscut to the footwall of 135 ft., or a distance at right angle to dip of 60 to 90 ft. Development work, reaching a depth of over 600 ft. on the dip of the vein, was done by means of various crosscut tunnels, winzes and laterals. A churn-drill hole was also sunk, which intersected the vein at a depth of 1100 ft. on the dip. Sufficient development work of this character was completed by 1937 to justify the erection of a reduction plant to roast the sulphide ore and cyanide the calcine, and to treat the oxidized ore by straight cyanidation, to make a doré bullion for shipment to the mint. The erection of the plant was started that year and finished early in 1938, with a combined capacity of about 800 tons per day. It has operated continuously since erection and recently was enlarged to handle 1000 tons per day. The vein dips with the slope of the mountain, and by stripping the hanging wall the ore may be mined by power shovel for several hundred feet below the outcrop. Thereafter underground methods will be necessary.

Jan 1, 1940

By F. J. Longworth

For a number of years an intensive study has been made to improve the blast-furnace practice at Copperhill not only as to costs but to provide a good grade of gas for the acid plants. This study, took us back to the fundamentals of successful pyrite smelting; namely, a blast furnace working on a clean sulfide ore, free from fines, and the use of quartz high in silica as a flux. In 1919, the quality of the flux was improved and its effect on the metallurgy was very marked. To improve the practice further, it was necessary to work out some method of treatment for the sulfide fines and the siliceous and refractory ores, other than in the blast furnace. A series of milling tests indicated that these products could be successfully treated by differential flotation, and in 1922 plans were made and a flotation plant built. It was estimated that the flotation plant would deliver to the smelter 40 to 50 tons of concentrates per day; the problem was what to do with these concentrates. It was out of the question to attempt to smelt them in a blast furnace because of the dust losses and, because our practice demands a charge free from fines for satisfactory operation. There is no custom smelter in the vicinity to which concentrates may be sold, pending the time when the tonnage will be sufficient to warrant roasters and a reverbatory furnace. Sintering and nodulizing were investigated but neither method fitted into our scheme of metallurgy. A close study of the converter practice was then made to see if some method could not be evolved to smelt this tonnage of concentrates in a converter without preliminary treatment. At that time, the converter installation consisted of two 12-ft. Great Falls converters, which were handling 12 to 14 per cent. matte. The heat balance of the converters working with 12 to 14 per cent. matte was more than sufficient to smelt the contemplated tonnage of concentrates but the mechanical difficulties. of feeding fines into the Great Falls converter made the use of this type of converter impractical.

Jan 9, 1924

By John E. Sherborne, William M. Newton

The value of the electric log as a means of interpreting underground structures has been increasingly demonstrated by its almost universal present-day use. It becomes important, therefore, to be able to evaluate the factors that affect the values obtained. By far the greatest number of electric logs are measured today in boreholes filled with rotary mud, the properties of which may exert an important influence on the electric log. In an effort to establish what the order of magnitude of such effects might be, a series of investigations was made in the laboratory and the field. Apparatus suitable for either field or laboratory use was designed and a procedure established for determining the factors that affect the resistivity of drilling muds over a wide variation in mud condition. Five muds representative of the types of clay-base muds employed in California were tested and the following observations were made: 1. The effect of raising the temperature from 8o° to i8o°F. is to decrease the resistivit~ of the mud or filtrate approximately 50 per cent. 2. The resistivity of the mud in most cases closely approximates that of its filtrate. 3. The change in the resistivities of muds caused by the addition of chemical is not the same function of the amount of chemical added for each mud. 4. The effect produced by increased sodium chloride content is to reduce markedly the resistivity of the mud and its filtrate. 5. Weighting materials, such as baroid and limestone, tend to increase the resistivity of drilling mud. 6. Cement and counteracting reagents reduce mud resistivity. The results of the investigation indicate that the mud resistivity is materially affected by the addition of electrolytes whether these are added for the purpose of conditioning the mud or enter the mud when salt beds or highly saline structures are penetrated. It should be observed, however, that the magnitude of the effect produced upon the electric log by changes in mud resistivity is also dependent upon the relative salinity of the water en. countered in the formation being logged. The laboratory data are supplemented by field data, which indicate that both the self-potential and resistivity curves are influenced by mud resistivity.

Jan 1, 1946

By Michael J. Messel

W. P. MOULD*—Has consideration been given to the problem of retreating the 4,000,000 ton tailings pile to recover the very considerable amount of asbestos fiber that was lost over the years due to insufficient operation and poor recoveries? Also, what was known about the possible content of chromite and the sulphides of nickel and cobalt? M. J. MESSEL (author&apos;s reply)—Consideration was given to the problem but nothing concrete has been done up to this time. W. P. MOULD—I am certain the tail- ings contain a very considerable quantity of fiber, not all of which is in the shortest lengths, because in the milling of talc rock from a neighboring location having a common origin with the asbestos bearing serpentine source of the tailings pile, a concentrate carrying 15 pct Ni and Co in complex sulphides and constituting about 4/10 of 1 pct of the talc rock was recovered. In the belief that the tailings also contained an appreciable content of chromite (then in demand, 1937) it was suggested to the management of the asbestos company that a truck load of the current tailings be sent to the flotation plant of the talc company where it would be put over the concentrating tables in an effort to discover the presence and roughly the percentages of chromite, nickel, and cobalt. Preparations for the test were woefully inadequate and the test was completely snarled up by the surprisingly large quantity of asbestos fiber from 1 in. long down to very short fiber, all of which matted up and plugged launders and pumps with disastrous effects and practically the total loss of the sample. If anything was indicaded by the test, it was that a wet process for milling asbestos was worthy of very serious consideration. M. J. MESSEL—A wet milling process for tailings retreatment only, is at present under study at our plant. One asbestos mining company in Canada has constructed a fairly large test plant for this purpose, but as yet no definite results have been obtained other than in a small test unit plant which were in some respects quite encouraging.

Jan 1, 1950

By Ming-Kao Yen

S. E. MADDIGAN*—I think Mr. Yen should be complimented on doing a very excellent and a very comprehensive piece of work. I was particularly pleased to note the use of X rays in a somewhat less tedious technique than pole figures. In the past I think many workers have been diverted away from the use of X rays in such investigations because of the large amount of work involved in pole figures if one is dealing with a great number of specimens. The same thing, I think, applies to the determination of orientations by counting of twins which is also a very tedious procedure; as a result several papers which have otherwise been of excellent quality have depended upon etching techniques for determination of orientation. The present paper indicates that the results reported in those previous papers were correct, but I think the use of etching techniques alone is a rather shaky basis for setting up theories in the literature. W. M. BALDWIN, JR.*—This discussion will be a short note of correction and apology. In my discussion of Brick, Martin and Angier&apos;s paper I incorrectly reported the orientation of phosphorus deoxidized copper annealed at temperatures. When I learned that Mr. Yen was preparing the present paper I asked that he correct this previous error, which he most graciously consented to do. However, through my ambiguity the correction noted on p. 59 of Mr. Yen&apos;s paper is still not the proper one. For this I take full responsibility. The orientation which I observed in phosphorus deoxidized copper at high temperatures is the (112) [111] which agrees with the pole figure given in Fig 8a, on p. 65 of Mr. Yen&apos;s paper. Our observations were made on the basis of twin counts rather than X ray determinations and therefore serve as an independent confirmation of Mr. Yen&apos;s data. M. K. YEN (author&apos;s reply)—The author wishes to thank Dr. Maddigan for his comments in regard to the application of glancing technique to textures. As a matter of fact, considerable work has been done recently in connection with

Jan 1, 1950

By I. J. Gardescu

The production of petroleum in Rumania in 1932 totaled 52,560,000 bbl., an increase of 6.8 per cent over 1931. The average daily production rose from 123,000 bbl. at the beginning of the year to 173,000 bbl. at the end of the year, the production in the Gura-Ocnitzei field rising from 38,800 to 80,700 bbl. per day. This increase was due to competitive development during which time the percentage of total production of the various operating companies remained practically unchanged. Chief producers were: Concordia, 30.3 per cent; Unirea, 21.6; Romano-Americana, 13.5; Steaua Romana, 8.3. Boldesti, where production of the previous year was practically doubled, would have produced far greater quantities were it not for the fact that both Astra Romana and Romano-Americana curtailed drilling operations. Of the total daily production of 21,800 bbl. a day, the Astra Romana, Steaua Romana and Romana Americana, produced respectively 67, 25 and 8 per cent. The Moreni field showed a sudden drop in production from the preceding year, due primarily to reduced drilling activities. Its production of only 21,000 bbl. a day removes it from first position for the first time in many years. All the other fields showed a decrease in production, except Ochiuri, 3 miles west of Gura-Ocnitzei, and Ceptura, 12 miles east of Boldesti. Drilling1 During the first nine months of 1932, 116 wells were completed, 8 of which were dry holes. The Meotian (lower Pliocene) is the producing horizon in 73, the Dacian (upper Pliocene) in 31, and the Oligocene in 4 of the completed wells. Three of the Oligocene wells were drilled in the Bacau district and one at Grausor, near Campina. In addition, 47 wells were drilling on Oct. 1,1932. Of the total completions and drilling wells, 92 per cent employed rotary for a footage of 98.5 per cent of the total drilled. Power for rotary drilling was furnished to 105 wells by steam, to 38 by electricity, and to one by gasoline.

Jan 1, 1933

HUBERT INGERSOLL ELLIS Hubert Ingersoll Ellis, who met accidental death on Jan. 6, 1919, in eastern Washington, had already advanced far in the profession of mining engineering and gave promise of a brilliant future. He was born on the Washington coast in June, 1889, and had his preparatory schooling at Menlo, Wash. In 1905; he entered the College of Mines of the University of Washington, and at once made a high record of scholarship. After completing 3 years of his technical course, he went to the interior of Alaska, where he spent 4 years in mining. During this time he was field correspondent for technical-magazines and wrote several important articles. In 1915, Ellis completed his senior year in the College of Mines; his graduation thesis was a comprehensive paper on mining methods at Fairbanks, Alaska, which was published as a series of articles in the Engineering and Mining Journal. On July 14, 1915, he entered the employ of the Bunker Hill & Sullivan Mining & Concentrating Co., Kellogg, Ida., as an underground miner, and was soon promoted to the position of operative on experimental metallurgy and later chemist in the same work. Later, he did exhaustive work on flotation with particular attention to the use of various addition agents. For the past 212 years, his work was largely in the field, examining mining prospects and directing work of investigation, during which time, while located at Sumter Valley, Ore., in 1916, he was elected to membership in the American Institute of Mining and Metallurgical Engineers. During the same year, he was married to Miss Grace Wells, of Arkadelphia, Ark., and besides his widow leaves a young son. In Hubert I. Ellis, there was found the rare combination of executive, administrative, and business ability, together with a temperament exceptionally well adapted and trained for research work and for the careful and exhaustive examination of engineering problems, both practical, and academic. These qualities were in addition to a splendid physique and a personality that made him highly successful in the handling of sub-ordinates and in his contact with his associates and the public.

Jan 4, 1919

By Roy A. Hardy

The Getchell mine is a comparatively recent discovery in the old Potosi mining district, Humboldt County, Nevada, a district organized in the seventies and eighties, in which some prospecting was done in several metamorphic zones along a monzonite-sedimentary contact. The district had no production in the early days and no further work was done until the same metamorphic zones were prospected for tungsten (scheelite) in 1917 and 1918, when a few small lots were shipped. It again lay dormant until 1934, when prospectors investigated a large siliceous escarpment at one end of the old district where outcroppings were a prominent feature of the topography. Undoubtedly, during the early prospecting of the district, the outcroppings had been investigated for mineral, probably by panning, with negligible results because the gold later found is invisible in the concentrate and is refractory. A fire assay revealed the presence of sufficient gold to warrant development work. The outcroppings were easily prospected by short tunnels having their portals in the siliceous material on the hanging-wall side, with a maximum distance to crosscut to the footwall of 135 ft., or a distance at right angle to dip of 60 to 90 ft. Development work, reaching a depth of over 600 ft. on the dip of the vein, was done by means of various crosscut tunnels, winzes and laterals. A churn-drill hole was also sunk, which intersected. the vein at a depth of 1100 ft. on the dip. Sufficient development work of this character was completed by 1937 to justify the erection of a reduction plant to roast the sulphide ore and cyanide the calcine, and to treat the oxidized ore by straight cyanidation, to make a doré bullion for shipment to the mint. The erection of the plant was started that year and finished early in 1938; with a combined capacity of about 800 tons per day. It has operated continuously since erection and recently was enlarged to handle 1000 tons per day. The vein dips with the slope of the mountain, and by stripping the hanging wall the ore may be mined by power shovel for several hundred feet below the outcrop. Thereafter underground methods will be necessary.

Jan 1, 1941

By S. A. Hoenig

In an earlier publication, the application of exoelectron technology to the monitoring of dry grinding in a ball mill was reported. Recently, the author has investigated the ball mill problem in more detail with the system shown in [Fig.1]. Here, some of the current generated in the ore-ball-water mixture was picked up by the metal chains and served as a measure of the rate of grinding. While the mechanism of current generation has not been completely clarified, a parallel study of electrical effects associated with rock drilling indicated that the major driving force is the zeta potential*, while oxidation of the freshly fractured material provides the remainder of the current. Both of these effects are sensitive to the exposed surface area, and it might be expected that the observed currents would provide a measure of the grinding rate. This "surface area" effect allows the measurement system to be used in any type of rock or ore grinding. Some studies with sulfide copper ore indicate higher currents than those observed with silica sand, but the overall pattern of current versus grinding rate is essentially the same. Experimental Studies The experimental system was set up as shown in [Fig. 1]. The initial studies were done with silica sand because of its availability and for comparison with the data of reference 1. The sand was sized to an average of 425 /MI (40 mesh); a typical charge consisted of 2.5 kg (5.5 lb) of sand, 0.85 L (0.22 gal) of tap water and 2.5 kg (5.5 lb) of iron grinding balls 25.5 mm (1 in.) in diameter. For a typical "run," the mill was loaded with a fresh charge of ore and operated for periods ranging from 10-60 min. During this time, the current was measured with a Keithley 417 picoammeter and a chart recorder. At the end of the grinding period, the charge was removed and sieved to recover the 425, 300, and 212 µm (40, 50, and 70 mesh) fractions which were then drained and weighed. A fresh charge was used for each test and every attempt was made to keep the draining and weighing processes uniform.

Jan 1, 1985