Search Documents

By George Holderer

THE relation between man-hours of labor and production may be correlated for any industry, and already it has been widely used in piece-work studies. It is not in general use as yet for recording labor efficiency of an enterprise or an industry. An exception to this is the Cement Institute, which tabulates the man-hours used in each operation and FIG. 1.-LABOR PRODUCTIVITY IN COPPER MINING. thus gets the total man-hours per barrel of cement. The writer knows of one steel company that keeps a complete record of man-hours for every operation from mining the ore to finishing the product. No doubt many others find the information useful, and such a study could be of value to the mineral industries.

Jan 1, 1935

By O. P. Quist, S. H. Carpenter

The internal friction of single-crystal LiF has been investigated as a function of crystal orientation, while simultaneously applying a static compressive load. Three different crystal orientations were used, these being samples with the (100), (110), and (111) directions along the sample length. Results show the internal friction measured to be similar in appearance but different in magnitude for the (100) and (110) samples. With both orientations, application of a static stress caused an increase in the amplitude-independent damping, which recovered with time. The amplitude-dependent damping was found to first decrease with load up to the yield stress and then increase rapidly. Samples of the (111) orientation were found to have no measurable internal friction even when loaded up to 3.0 kg per sq mm. These results are discussed in terms of current theories of internal friction in ionic crystals. This paper presents the results of an investigation of the dislocation mobility and dislocation motion in high-purity single-crystal LiF at applied stress levels up to the yield stress. Internal friction measurements of the dislocation damping while simultaneously applying a static compressive load was the experimental technique used. LiF crystals much prefer to slip on the (110) set of slip planes. However, with a suitable stress distribution they can be forced to slip on the {100} set of slip planes.' For both sets of slip planes the slip direction is the same, namely (110). The object of this investigation was to try and learn more about the dynamical properties of dislocations on the different slip systems in LiF crystals. EXPERIMENTAL PROCEDURE Dislocation damping was measured, in terms of the log decrement, using a five-component marx2 composite piezoelectric oscillator. The apparatus used allows one to measure the dislocation damping while simultaneously applying a static compressive load on the sample. A schematic of the apparatus used is shown in Fig. 1 and a complete discussion of its construction and operation is given elsewhere.3 All measurements were made at room temperature with a longitudinal resonant frequency of approximately 50 kc per sec. Damping measurements were made using longitudinal vibrations with strain amplitudes from up to 10. Static loads up to 3.0 kg per sq mm were applied to the sample while the damping was being measured. Samples used in this investigation were high-purity single crystals of LiF purchased from the Harshaw Chemical Co. All samples were approximately of an in. square and were hand-ground in length to one-half wavelength to match the resonant frequency of the driver and gage quartz crystals. Samples of three different crystallographic orientations were used, namely, crystals with the (loo), (110), and (111) direction parallel to their length. The magnitude of the damping is a direct measure of the area on the slip planes swept out by dislocations during a complete oscillatory cycle. With a sample of the (100) orientation the maximum applied shear stress, due to longitudinal vibrations, is on the (110) slip planes with zero shear stress on the (100) slip planes. Thus, damping measured from a (100) sample is due entirely to dislocation motion taking place on the (110) set of slip planes. Similarly, for a sample of the (111) orientation there is zero shear stress on the (110) set of slip planes and the damping measured from a sample of this orientation is due entirely to dislocation motion on the (100) set of slip planes. Samples of the (110) orientation will have a component of shear stress on both sets of slip planes and

Jan 1, 1969

G. D. Kneip, Jr., and J. 0. Betterton, Jr. (Union Carbide & Carbon Corp., Oak Ridge, Tenn.)—The authors have contributed to the theory of zone melting by considering the effects of the solidification of the final zone on the distribution curves for the finite length bar. However, they do not consider in sufficient detail the restrictions imposed by the particular metals involved. Eq. 4 has the analytical solution which is equivalent to Pfann's' equation for normal solidification of the final zone. For distribution coefficients less than one, it is apparent from this form of the equation that, as the solidification approaches completion, the concentration attains extremely large values which are inconsistent with the density of physical materials. The necessary restrictions then are that the solute concentration cannot exceed the density of the alloy, or more frequently, the solubility limit in the solid phase. Secondly, the distribution coefficient cannot be constant from 0 to 100 pet solute unless the liquidus and solidus coincide over this whole region. It would thus be more correct to impose limits on the solubility such as would be indicated by a typical eutectic diagram. Furthermore, in this case the assumption that the distribution coefficient remains constant is more likely to be realized. Eq. 4 should then be replaced by the following C,,(x) =k/L-x [ ?11 3 C11 (x)Ddx - ?11 4 Cn(x)dx= k[Cn (L-l)/k] [ L-x/l]x-1 L-l = x = L-l [kCenterlie/Cn (L-l)]1/k-1 The typical shape for the distribution curve in the final zone is indicated schematically in Fig. 12 for the simple eutectic case. The solidification proceeds according to Pfann's' expression until the concentration reaches the maximum solubility in the solid phasc. At this point, the concentration changes abruptly to the eutectic composition and remains at this value for the duration of the solidification. The width of the flat region, and the back reflection of this effect into the concentration curve on subsequent passes, depends upon the ratio of the eutectic concentration to the original concentration, upon the distribution coefficient, and upon the number of passes. Similar effects on the concentration curves in the first zone length would be expected in many systems for which the distribution coefficient is greater than one, and where the maximum solubility is not too much larger than the original concentration. The discussers agree with the authors that the assumption that the liquid is uniform in concentration should be considered with caution. Kneip and Better-ton' have shown, however, that in floating zone refining of zirconium, using induction heating, the distribution of iron agrees with theory which assumes a uniform liquid composition. Hence, in this case, the experimentally realized distribution coefficient agrees with the present phase diagram within the experimental uncertainty. L. Burris, Jr., C. H. Stockman, and I. G. Dillon (authors' reply)—The authors are happy to have received the comments of G. D. Kneip, Jr., and J. 0. Betterton, Jr., which provide additional insight into the zone refining process. It is true that the equations developed are inapplicable if the solute solubility in the solid phase is exceeded. However, this restriction on the use of the equations was clearly stated in the paper. A closer approach to

Jan 1, 1957

By James B. Graham, Earle E. Smith

The value of all minerals produced in the United States for the years 1930 to 1932 averaged $3,460,000,000 annually, according to Bureau of Mines reports. Of this amount, approximately $1,280,000,000 represents nonmetallic substances, such as petroleum, gas, sulfur and salt produced as fluids. Thus, about.37 per cent of the country's mineral wealth is procured through wells drilled in the earth, which require casing to keep them open throughout their life. Deep well drilling has increased greatly in the last few years. Several strings of casing have been set deeper than 9000 ft.—one, a string of 7?-in. O.D. casing, 10,301 ft. long, run in a well drilled by the Gulf Production Co. in West Texas, holds the record up to this time. These are longer strings than were contemplated when the American Petroleum Institute Specifications were adopted 10 years ago, therefore it is necessary to consider whether the present joint standards are adequate for such strings, or for even longer strings that may be called for in the future. Petroleum engineers are asking: Is there a possibility that stronger joints can be designed without going to stronger and more costly steels than offered in present specifications? Without sacrificing too greatly the tensile properties of the joint, can its design be changed to allow more clearance in the open hole below the next larger string of casing, and thus avoid costly underreaming? What can manufacturers offer for drilling through heaving shale? Fundamentals of the Problem During the last few years considerable study has been given to casing joints and their limitations, and since attention must be given to this subject by those who will be required to drill deeper holes, it is timely to review the fundamentals of the problem. These will be recognized generally as the following: 1. No threaded joint can be stronger than the area of metal under the root of the first perfect thread, and this area must be large enough

Jan 1, 1935

By J. N. Anderson

Developments in reverberatory furnace practice at Noranda over the period 1928 to 1953 are described. Features of interest are increasing furnace tonnage from 700 to 2000 tons per furnace day, the use of the suspended basic roof, and improvements in furnace fuel ratio. Points of comparison and contrast between Canadian and American furnace practice are given. OPERATIONS at the Noranda smelter have been described in a series of papers published between 1930 and 1945.1-4 In the present paper the development of reverberatory furnace practice over the 25 year period from the beginning of operations up to the present time will be reviewed and current furnace practice will be described in some detail. Particular attention will be given to those features which are characteristic of Canadian practice: first, the use of the suspended basic roof; second, the deliberate driving of furnaces to give tonnages in the range of 1400 to 2000 tons of solid charge per furnace day; and third, the continuous operation of furnaces at high tonnage for many years without any extended shutdown for repairs or rebuilding. Finally, points of comparison and contrast between reverberatory furnace practice in Canada and the United States will be discussed. The Noranda smelter, which started in December 1927, was built for the reduction of ore from the Horne mine where early development had shown a comparatively small tonnage of rich cop-per-gold ore of a grade and character which would require direct smelting. The smelter was designed for a capacity of 1000 tons of ore a day and originally consisted of eight roasters, two reverberatory furnaces, and two converters, together with necessary auxiliaries. The discovery of a large new body of ore in 1928 made it necessary to increase both the size and the scope of the reduction plant. A mill which was ultimately brought to a capacity of 3000 tons a day was built for concentrating low grade ore and in addition the capacity of the smelting plant had to be greatly expanded, so that for the next fifteen years every effort in the smelter was directed toward achieving maximum tonnage. Because the layout and design of the smelter made it impracticable to add a third furnace, it was necessary to increase the tonnage on the two original furnaces by improvements in furnace design and operation, roasting and converting equipment being added as required to meet the demands of increased furnace capacity. The smelter tonnage, which reached a peak in 1941, was curtailed in 1944 when it became necessary to operate on a one-furnace basis because of a reduction in the output of the Horne mine due to a shortage of labor. Two-furnace operation at a moderate tonnage was resumed in 1949. In the early years the smelter tonnage was made up almost entirely of Noranda ores and concentrates. However, since 1937 an increasing tonnage of ores and concentrates from other mines has been treated in the smelter. In recent years, with decreased production from the Horne mine, copper concentrates from several mines in the district, treated on a custom basis, have formed an important part of the smelter tonnage. A review of developments in reverberatory furnace practice at Noranda falls naturally into the following four periods: 1—1928 to 1935, the early years; 2—1935 to 1944, the period of greatest change and development; 3—1944 to 1949, one-furnace op-

Jan 1, 1955

By H. F. Yancey, M. R. Geer, C. L. Allyn, R. H. Eckhouse

Four coals were treated in the Humphreys spiral concentrator, and the products were examined by float-and-sink and screen-sizing tests to determine fundamental performance characteristics. The efficiency of the separation between coal and impurity was shown to be influenced greatly by particle size. FOUR coals exhibiting different washability characteristics were washed in the Humphreys spiral concentrator and the products examined by float-and-sink and screen-sizing tests to determine the mechanics of the separation between coal and impurity. With one coal, each 1/2-in. increment of the full width of the stream in the spiral was examined separately to determine where each size and specific-gravity fraction of the feed was stratified. The basic performance characteristics of the spiral were found to be the same for all of these coals. The coarsest fraction of heavy impurity stratified so far out in the stream that it could not be removed through the refuse ports and thus either entered a middling product or contaminated the washed coal in the case of a two-product separation. Impurity particles finer than about 100-mesh also were carried out in the main body of the stream and therefore were not removed in the refuse product. Little loss of clean coal in the refuse occurred in sizes coarser than 28-mesh or finer than 100-mesh, but considerable coal of intermediate size stratified in the stream that was drawn off through the refuse ports. Because of this modifying influence of particle size, the spiral is unable to make an efficient two-product separation between coal and impurity without some retreatment. Recirculation of a middling product through the same spiral as a means of obtaining more efficient operation was not attempted in this investigation. The size and specific-gravity composition of the middling product are, however, such as to render it amenable to simple hydraulic classification for recovery of coal. Retreatment of a middling product would improve the efficiency of the spiral, but even when making a substantial amount of middling, considerable clean coal enters the refuse product. Consequently, retreatment of a combined refuse-middling product would appear to offer greater promise for providing maximum efficiency. Because of the sizing characteristics of the spiral, a classified feed can be treated with higher efficiency than is possible on a natural raw coal. For this reason, semiclassified feeds, such as silt-bank or classifier-underflow materials, can doubtless be treated with higher efficiencies than those shown for raw coals in this report. Performance is, of course, only one of many factors that enter into the choice of cleaning equipment. The fact that the one company now using the spiral for cleaning coal has built a second spiral plant is ample evidence that the inherently low efficiency of this unit may be overshadowed by the low operating costs afforded by its extreme simplicity. Introduction A number of factors enter into the selection of coal-cleaning equipment, and most of them bear directly on the final cost of prepared coal. One of these factors is performance of the cleaning unit in terms of the efficiency of the separation it effects between clean coal and refuse. The importance of performance in the selection of a coal-cleaning unit varies with the difficulty of the cleaning problem. Obviously, if the character of the coal is such that only certain types of cleaning units are capable of yielding a clean product meeting market requirements, performance is of prime importance. If, on the other hand, the cleaning problem is so simple that any type of cleaner will provide coal of suitable quality, performance is important only insofar as it affects costs by determining the amount of salable coal lost in the washery refuse. With the objective of providing the coal industry with information that is helpful in the selection of

Jan 1, 1951

By R. H. Eckhouse, M. R. Geer, H. F. Yancey, C. L. Allyn

Four coals were treated in the Humphreys spiral concentrator, and the products were examined by float-and-sink and screen-sizing tests to determine fundamental performance characteristics. The efficiency of the separation between coal and impurity was shown to be influenced greatly by particle size. FOUR coals exhibiting different washability characteristics were washed in the Humphreys spiral concentrator and the products examined by float-and-sink and screen-sizing tests to determine the mechanics of the separation between coal and impurity. With one coal, each 1/2-in. increment of the full width of the stream in the spiral was examined separately to determine where each size and specific-gravity fraction of the feed was stratified. The basic performance characteristics of the spiral were found to be the same for all of these coals. The coarsest fraction of heavy impurity stratified so far out in the stream that it could not be removed through the refuse ports and thus either entered a middling product or contaminated the washed coal in the case of a two-product separation. Impurity particles finer than about 100-mesh also were carried out in the main body of the stream and therefore were not removed in the refuse product. Little loss of clean coal in the refuse occurred in sizes coarser than 28-mesh or finer than 100-mesh, but considerable coal of intermediate size stratified in the stream that was drawn off through the refuse ports. Because of this modifying influence of particle size, the spiral is unable to make an efficient two-product separation between coal and impurity without some retreatment. Recirculation of a middling product through the same spiral as a means of obtaining more efficient operation was not attempted in this investigation. The size and specific-gravity composition of the middling product are, however, such as to render it amenable to simple hydraulic classification for recovery of coal. Retreatment of a middling product would improve the efficiency of the spiral, but even when making a substantial amount of middling, considerable clean coal enters the refuse product. Consequently, retreatment of a combined refuse-middling product would appear to offer greater promise for providing maximum efficiency. Because of the sizing characteristics of the spiral, a classified feed can be treated with higher efficiency than is possible on a natural raw coal. For this reason, semiclassified feeds, such as silt-bank or classifier-underflow materials, can doubtless be treated with higher efficiencies than those shown for raw coals in this report. Performance is, of course, only one of many factors that enter into the choice of cleaning equipment. The fact that the one company now using the spiral for cleaning coal has built a second spiral plant is ample evidence that the inherently low efficiency of this unit may be overshadowed by the low operating costs afforded by its extreme simplicity. Introduction A number of factors enter into the selection of coal-cleaning equipment, and most of them bear directly on the final cost of prepared coal. One of these factors is performance of the cleaning unit in terms of the efficiency of the separation it effects between clean coal and refuse. The importance of performance in the selection of a coal-cleaning unit varies with the difficulty of the cleaning problem. Obviously, if the character of the coal is such that only certain types of cleaning units are capable of yielding a clean product meeting market requirements, performance is of prime importance. If, on the other hand, the cleaning problem is so simple that any type of cleaner will provide coal of suitable quality, performance is important only insofar as it affects costs by determining the amount of salable coal lost in the washery refuse. With the objective of providing the coal industry with information that is helpful in the selection of

Jan 1, 1951

By C. A. Allen

FOR over three years we have been endeavoring to reduce the number of fatalities and minor accidents in Utah. The physical condition of the o? Represents the average of 4 States mentioned excluding the catastrophes indicated. FIG. 1.-NUMBER KILLED PER MILLION SHORT TONS OF COAL PRODUCED IN COLO¬RADO, NEW MEXICO, UTAH, WYOMING, AND BITUMINOUS MINES OF THE UNITED STATES, 1914 TO 1920. mines, concentrators, and smelters is good, yet the number of accidents,

Jan 5, 1922

Development: Iron Ore In Peru Utah Construction Co., of San Francisco and a Peruvian Government corporation have signed an agreement for development of a large iron orebody in Peru. Most tests show 60 pct iron or better. Most of the ore is magnetite. Although ore will be shipped mainly to Eastern U. S. blast furnaces, shipments to the West Coast may develop as the orebody opens up. The construction firm indicated three American steel companies have shown interest in the project. The orebody is located 300 miles south of Lima and about 15 miles from the coast at the same longitude as New York City. Allen D. Christensen, executive vice-president of Utah Construction noted that the distance is shorter between Peru and the East Coast. The ore deposit measures about four to five miles in one direction and 10 miles in the other. As many as 15 outcroppings within these limits have been reported. Several of the outcroppings are thousands of feet long by several hundred feet wide, Christensen said. Shipments between 100,000 and 200,000 tons of ore per month are projected by the company. Cost of opening the mine to production may go over $5 million, Christensen said. It would include building of docks for ocean going vessels, housing, and heavy duty mining and processing equipment. The contract with the Peruvian government gives Utah Construction six to eight months for additional tests and a 20 year concession for mining the orebody. Peru will collect royalties on all ore. The company will also supply about 300,000 tons per year for an electric furnace to be installed by the Peruvian Government at Chimbote. No steel is produced in Peru now. Utah Construction formed the Argonaut Mining Co. last year, and was shipping iron ore from British Columbia to Japan at a rate of more than 50,000 tons a month, before winter weather decreased operations.

Jan 1, 1952

By A. Blake Caldwell

Northwest frontier is here defined as Alaska, the Yukon, the Northwest Territories and northern British Columbia. The mineral potential of this area is great and will unfold in the decade ahead but now the largest single producer is the Pine Point lead-zinc mine in the NWT about 1200 km north of Edmonton, Alta., close to the south shore of Great Slave Lake. The mill, initially designed for 4540 mtpd, was expanded to 7250 mtpd in 1969. This operation is described in detail in Mining and Concentrating of Lead and Zinc, Vol.1, Chapt. 17, SME/AIME in the article by W. H. R. Gibney.

Jan 1, 1971

By E. M. Frizzell

Introduction In 1956, the first mobile crusher was installed in a limestone quarry in Hover, Germany. These early mobile crushers solved the problem of wet and soft ground conditions that did not permit the use of haulage trucks due to the high cost of building and maintaining the haulage roads. Quarry operators also wanted to take advantage of continuous belt conveyor haulage systems and the resulting cost savings. European mines, particularly the coalfields, are in soft materials that allow use of bucket wheel excavators and their following continuous haulage system of belt conveyors. Quarry operators, with this background knowledge of belt conveyors in mining, easily accepted the conveyor concept, once in-pit mobile crushing solved the problem of size reduction of mine-run-rock required for conveyor application. Four years after the initial mobile crusher installation, the first report published on the subject appeared in the December 1960 issue of Mining Engineer (Kochanowsky, 1960). With this early background and basis for direction, mobile in-pit crushing and conveying have increased to more than 74 installations. The capacity, type of crusher, and material being processed appear in Table 1. These data in Table 1 indicate that experience to data has been heavily influenced by limestone installations of less than 1 kt/h (1102 stph). Most of these installations were directly fed by a mine face excavator because of the limestone quarry ground conditions. Crusher Configurations Crushers systems developed to date have varying mobility capabilities. These range from fully mobile units that are used continuously in this mode to fixed, permanently located in-pit crushers. The following terms are presented to help distinguish the range of mobility within the configurations of "mobile" crusher systems planned or now in use. Mobile Crusher The crusher works at the mine face, is directly fed by an excavator, and moves in unison with the excavator on its own transport mechanism as the mining progresses. An example is the unit delivered in 1976 to Yamama Cement Co. in Reyadh, Saudi Arabia.

Jan 1, 1986

By Walter S. Weeks

THE subject of the flow of air through a mine under the influence of natural ventilation has proved an elusive one. To my knowledge only cut and try algebraic solutions have been used where multiple paths are involved. The object of this paper is to present a method for the graphical determination of the flow induced by the natural ventilation pressure (N.V.P.) alone or in combination with a fan. The N.V.P., measured in pounds per square foot, induced by two columns of air between two horizontal planes at different elevations, is equal to the difference in weight of the two columns, assuming each to have a cross-sectional area of 1 sq. ft. Pressure in pounds per square foot is converted into inches of water by dividing by 5.19. Methods of calculating the N.V.P. have been discussed in the literature. In this paper, for simplification of exposition, the change in pressure due to the weight of superimposed air will be assumed directly proportional to the length of the column. The actual methods of calculating the N.V.P. do not affect the solutions to be presented. It must be realized that the precision to be obtained from the small- scale, inked diagrams presented is not as great as would be obtained from large-scale, penciled curves that would be used in practice. The accuracy of the curves does not warrant reading "volumes" closer than the nearest thousand cubic feet per minute. CASE I. DUCTS IN SERIES The simplest type of problem is to determine the flow when two shafts are connected by a crosscut, as shown in Fig. 1. In this example, the resistance of the shafts is to be taken as 0.0001 in. of water for a flow of 1000 cu. ft. per minute for a length of 1000 ft. The resistance of a crosscut is to be taken as 0.0005 in. of water for a flow of 1000 cu. ft. per minute. The small pressures necessary to cause changes in velocity will not be considered, because the accuracy of other obtainable data does not warrant their consideration.

Jan 1, 1938

By G. R. St. Pierre, M. J. Pool, R. Speiser

The activities of chromium in solid Cr-Ti solulions contaitning from 10 to 90 at, pct Cr were measured over the temperatutre range 1250" to 1380°C. The Knzudsen effusion technique with direct weighing of the Knudsen cell was used. This was possible because the vapor pressure 0f chromium is about 600 times greater than the capor pressure of titanium so that the amount 0f titanium effusing could be neg-Lected. the activities of titanium were determined by g-raphical integration of the Gibbs-Duhertz equatiotz. The solutio~ls sllcdied exhibited positive deviations from Raoult's law, but they approached ideality as the temperature was raised from 125 to 138t C. The integral free energy of mixing was calclated from the activity values. AF has a maximum value of - 1.66 kcal a1 1390°C and a composition oj 55 at. pcl Cv. The free energy of formation of 'TiCr, is —0.76' lzcal a1 1250°C. The heats and entropies of mixing usere determined using the second law and thus tile mlues could be in error by a considerable amount. The Cv-Ti system is not a regular solution. ThE increasing emphasis on application of various models to alloy formation and solution behavior has led to an intensified interest in the thermochemistry of alloys. In a majority of the cases where different models have been applied the alloy system used was a simple one. However, the amount of data available on more complicated systems is small and in order to fully understand the behavior of metallic solutions more experimental work has to be carried out on these systems. The Cr-Ti system is reasonably complex and even though this system was being investigated as early as 1937 by Kroll,' no previous work has been done on the thermodynamic properties of either the liquid or solid state. This research was undertaken to determine the thermodynamic properties of solid Cr-Ti alloys. EXPERIMENTAL PROCEDURE The experiments on the Cr-Ti system were carried out in a water-cooled metal vacuum system. This system is composed of a brass vaporization chamber wound with copper tubing for cooling purposes, a brass liquid-nitrogen trap, an oil diffusion pump, and a Duo-Seal forepump. The pressure in the system is measured with an ionization gage. The interior of the vaporization chamber is shown schematically in Fig. 1. The Knudsen cell is supported by the molybdenum tripod, and a magnetically operated shutter is provided to protect the Pyrex sighting glass against deposition of metallic vapors in the intervals between temperature measurements. Heating is accomplished by means of a 20-kw 450-kc Westinghouse oscillator. The susceptor in this case is 10-mil tantalum seamless tubing which also forms the outside of the sample container. The remainder of the sample container consists Of an im~ervims re-crystallized alumina crucible and a 10-mil-thick tantalum lid. Two holes are drilled in the lid. The diameters of these holes were measured with the calibrated stage of a "Tukon" microhardness tester. The thicknesses were measured with a micrometer. A Leeds and Northrup disappear ing-filament optical pyrometer was used to measure the temperature. This is sighted through the smaller of the two holes in the lid and corrections for the emissivity of the sample were not necessary since the ratio bf cell length to hole diameter was large enough to insure a hohlraum. The pyrometer was calibrated at the ~~t~~~l Bureau of Standards, and the correction for the Pyrex sighting window was determined in the laboratory. The titanium used in this investigation was obtained from Mallory-Sharon and the chromium from The Union Carbide Metals Co. The reported purities were

Jan 1, 1968

Two career employees of the US Dept, of the Interior, in a recent study, indicated that by the end of 1974, 73% of all public lands had been entirely or partially closed to exploration and development of essential resources for leasable minerals and 67% had been similarly restricted for locatable minerals.1 This has come about because of the lack of coordination within the executive branch and between the executive and legislative branches of the federal government and without regard for the cumulative, depressing effect on future domestic mineral production. Much of the withdrawn land is felt to have great potential for discovery of significant mineral deposits.

Jan 5, 1976

By C. A. Zapffe

BOTH the foundryman and the theoretical metallurgist are now generally agreed that the anomalous "rising" or "bleeding" of certain ferrous castings of killed metal is primarily attributable to hydrogen. How and why the phenomenon occurs, however, have not been made clear. The following discussion is intended to harmonize foundry experience with metallurgical theory so that the foundryman may possibly improve his ability to prognosticate whether a certain casting may bleed or not. BLEEDING, RISING, AND PIPING The "rising" to be discussed here is not to be confused with the action in rimming ingots as a result of the carbon-oxygen reaction, although a conclusion will be drawn that may apply to the art of controlling the rimming action. Only killed metal is considered, which includes both steels and irons. In Fig. I, three killed, chill-cast ingots are shown: a porous ingot whose evolution of gas during late stages of solidification has ejected metal through the top; an ingot in which moderate gas evolution during solidification has lifted the mushy central metal, i.e., it has "risen"; and, third, a sound ingot with good piping characteristics. The first type is referred to as "bleeding," and generally infers that a solid skin had first formed through which the liquid metal later erupted. There is no fundamental distinction between bleeding and rising in reference to killed metal.* POROSITY Hydrogen appears as a cause of bleeding because the gas evolution simultaneously leaves a porosity of a type traceable to hydrogen. Twenty years ago, Melmoth1 noted that killed steel that rose unexpectedly in the mold, if returned while liquid to the furnace and heated further under a reducing lime slag, would again rise in the mold; but that if the ingot were allowed to solidify and was then remelted no rising occurred. A critical quantity of the responsible gas is therefore removed by the rising action. Melmoth associated rising with high silicon (0.7 to 0.9 per cent) and concluded that silicon increased gas absorption. Because silicon is a strong deoxidizer, oxygen must, therefore, be eliminated from consideration of the three common gases in steel-oxygen, nitrogen, and hydrogen. Butterworth2 later decided that pinhole porosity was caused by gases dissolved in the metal and that hydrogen was obtained from steam in the mold. Allen's classical work on copper castings,3 soon followed by

Jan 1, 1942

By B. B. Zavoico

The oil industry in the Soviet Union closed the 1929-1930 operating year fulfilling its assigned program. During this period of time, however, no basic improvements were noted within the industry. A careful study of the conditions indicates that disorganization in most branches increased continually; that supply and transportation problems are approaching a state of breakdown; that labor problems become increasingly difficult and that the estimates for 1931-1933 period are too large. The production of all Russian fields increased by 27,200,000 bbl., as compared with the 1928-1929 operating year, while the next year's program anticipates a further increase of 46,000,000 bbl., and the 1933 plans call for an amazing total of 330,000,000 bbl., or over 150 per cent, more than the production of the operating year that has just closed. Table 1 shows petroleum production in Russia by individual areas and fields from 1916, and furnishes estimates for the three coming years, which constitute the balance of the Five-year plan. A brief analysis of Table 1 will show that the Soviet Planning Commission anticipates that the fully developed fields of the Apsheron-Baku will double their production by 1933, .as compared with 1929-1930, and that the old Grozny fields, which are already showing a tendency to decline, will increase their output from 44,110,000 bbl. in 1929-1930 to 68,000,000 bbl. in 1933. Furthermore, in 1933 new fields are slated to produce not less than 100,000,000 bbl., a figure which even under most favorable conditions in Russia would be hard to achieve. Of course, Russia could develop such production, because the country is not lacking in the prospective reserves, but the transportation in the country today (as in the past.) is in such poor shape that neither the supply of a gigantic oil industry nor the movement of its products is within the realm of possibility. Table 1 shows the future estimated crude oil production as required by the Five-year plan, and also as reasonably estimated by the author, taking into account natural declines in old fields, slow development of new areas and the deficiency of the transport. While in the 1929-1930 operating year only two major fields underproduced (Surakany and Maikop) and the plan was fulfilled to the extent of 97.4 per cent., the production of the last three months of 1930 ("Shock Quarter") indicated

Jan 1, 1931

By Francois E. Heuze

Today, the favored option for disposal of high-level nuclear wastes is their burial in mined caverns. As an alternative, the concept of deep disposal by rock melting (DRM) also has received some attention. DRM entails the injection of waste, in a cavity or borehole, 2 to 3 kilometers down in the earth crust. Granitic rocks are the prime candidate medium. The high thermal loading initially will melt the rock surrounding the waste. Following resolidification, a rock/waste matrix is formed, which should provide isolation for many years. The complex thermal, mechanical, and hydraulic aspects of DRM can be studied best by means of numerical models. The models must accommo¬date the coupling of the physical processes involved, and the temper¬ature dependency of the granite properties, some of which are subject to abrupt discontinuities, during [a]- ß phase transition and melting. This paper outlines a strategy for such complex modeling.

Jan 1, 1982

By G. M. Roberts

During the six years that the writer was connected, as chemist and chief assayer, with the Proprietary Mines, Broken Hill, N. S. W., several interesting experiments were made in the sampling of the silver-lead bullion obtained from the waterjacket smelting-furnace.

Jan 1, 1899

By R. I. Jaffee, H. R. Ogden, D. J. Maykuth, W. L. Finlay

Titanium with up to 7.5 pct Al forms single-phase a alloys that are work hardening, not heat treatable, and ductile as welded. The high aluminum Y phase alloys are not usefully ductile, despite their low hardness, but they have excellent oxidation resistance and hot hardness. LITTLE has been published on the effect of aluminum on the mechanical properties of titanium, despite the fact that two of the first four commercial alloys released by titanium alloy producers contained aluminum as an important alloying ingredient. These were both a-ß type alloys, one containing 4 pct A1 and 4 pct Mn and the other containing 3 pct A1 and 5 pct Cr. However, neither of these alloys reflects the true alloying nature of aluminum, since each contains important amounts of the ß-stabilizing addition, manganese or chromium. In a previous publication on the constitution of Ti-A1 alloys,' the authors showed that aluminum had a high solubility in a titanium and was strongly a stabilizing in nature. About 25 wt pct A1 is soluble in a titanium at room temperature. With as little as 5 pct added aluminum, the a transus temperature is increased from 882" to 970°C, and the correspond- ing ß transus to 995°C. The first intermediate phase in the Ti-A1 system is the ordered face-centered tetragonal TiAl phase, which extends from 34 to 46 wt pct Al. The present paper presents hardness data over the entire titanium-rich region of the system, through the TiAl phase, and mechanical property data on the alloy range capable of being fabricated. Methods of Investigation The materials used and general preparation of alloys, as well as experimental techniques used in the investigation of Ti-A1 alloys, have been adequately described.' Briefly, the preparation consisted of arc-melting the desired amount of high purity aluminum (99.99 pct purity) and iodide titanium (maximum hardness 100 Vickers, as deposited) together on a water-cooled copper hearth under a partial pressure of argon. The resulting button ingot was of such form that it could be fabricated without further treatment. All work was done using ingots weighing 10 to 15 g, which were hot-rolled to strip form whenever possible. Tensile samples were 3.25 in. long, 0.5 in. wide, and 0.040 in. thick, with a reduced section 1.25 in. long by 0.250 in. wide. The gage length used was 1 in. Electrical strain gages mounted on the

Jan 1, 1954

Upon nomination of the President, _the following Committee on Membership was appointed: Karl Eilers, Chairman, Arthur. S. Dwight, Lewis W. Francis, Louis D. Huntoon, Arthur L. Walker. The following resolution was then unanimously passed: WHEREAS, the integrity and life of this nation is dependent upon its ability to successfully defend itself against foreign aggression and attack, and WHEREAS, it is necessary that the nation he properly prepared to the end that our untrained patriotic volunteers he not uselessly slaughtered and our country disgraced and defeated, be it therefore RESOLVED: That the Board of Directors of the American Institute of Mining Engineers believe that compulsory universal military training should be at once instituted and that adequate, complete and immediate preparedness, should he provided. BE IT FURTHER RESOLVED: That, we believe that thorough and complete preparedness in the army, the navy and the industries of the country is vital; anything less is unsatisfactory and inefficient and may result in disaster to the nation.

Jan 5, 1917