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By Ralph Arnold

WHILE much geological work and drilling have been done in Venezuela, the incompleteness of geological evidence obtainable and the restricted areas in which drilling has been done make any estimates of oil resources extremely speculative. Nevertheless, it is felt that the data given are justified, in the belief that the ideas of a geologist familiar with the region are better than those of a promoter. The oldest beds of interest to oil men are those of the Cretaceous, which are well developed in eastern Venezuela and about parts of Lake Maracaibo. In both cases, a hard massive limestone is overlain by a great thickness, most of which is shale. In the west, in places this shale is more than 2000 ft. (610 m.) thick, while in the east it may attain 3500 ft., (1067 m.). The limestone and parts of the shale are oil producing. In the east, the Tertiary where present is almost entirely covered by the Quarternary, but in the west it is well developed. In probable unconformity over the Cretaceous shales is found a series of sandstones and shales, which may beg called the lower coal series. It is well developed southwest of the lake, where it is the oil container, and is probably the lowest series exposed east of the lake. Above this is the Pauji shale, which is probably the point of origin of all the oil east of Lake Maracaibo and in the State of Falcon. It has a thickness of about 3000 ft. (914 m.) in the Mene Grande region. The upper coal series is developed as a great thickness of sands, shales, and clays, with some partly consolidated limestone. Its thickness may be as great as 10,000 ft. It is the oil container of the Falcon-El Mene region and a probable location for oil in some of the area about the lake. It may or may not include the so-called Maracaibo series in which the oil is found at Mene Grande. Over much of Venenzuela, the Quarternary is present in unknown thickness, which probably is not great. It obscures oil structure in all cases and presents the greatest hazard of oil exploration for the fields which probably exist below.

Jan 7, 1922

By C. W. Wark, Fred Wise

The Getchell mine, a comparatively recent gold discovery, is in the old Potosi mining district, Humboldt County, Nevada. All ore is mined from open pits using Diesel shovels and gasoline trucks. Two types of ore are mined, oxide and sulphide. Oxide ores are siliceous or shaly; sulphide ores either gougy, siliceous or shaly. In mining, the sulphide ore is segregated into hard and gougy ore, because the gouge must be dried before it can be crushed. The oxide ore is not segregated. General Metallurgy The oxide ore is the leached outcrop that overlies the sulphide ore body. All of the iron and arsenic sulphides found in the sulphide ores have been thoroughly oxidized. The sulphur and arsenic have been leached out while the iron is present as hematite and limonite. The oxide ore is amenable to direct cyanidation, and although no visible free gold is present it is released at a comparatively coarse grind. The sulphide ore is more refractory, since the gold is all of micron size and some of it is locked in fine sulphides. It has been established microscopically that the free gold is coated with an impervious shell, which is probably some form of iron compound. The sulphide ores contain from 1.5 to 2 per cent of arsenic, which is present as orpiment and realgar with minor amounts of arsenopyrite. The orpiment and realgar are of a later mineralization than the gold and are barren. Roasting and washing of the sulphide ore prior to cyanidation is necessary to eliminate , the arsenic, to alter the gold coating from an insoluble form to a soluble form, and to release the gold associated with the fine sulphides. Washing the ground calcine in weak sulphuric acid, trona, or hot water are beneficial in the order named, but to date the water is the best from an economic standpoint. Metallurgical tests as well as plant practice indicate that a certain amount of gold cannot be recovered even by going beyond economic practices. On this basis metallurgical results are viewed in the light of an optimum tailing rather than a percentage of extraction. Milling Practice Crushing (Fig. I).—There are three primary ore bins, one each for oxide ore, gougy sulphide ore, and hard sulphide ore. In front of each bin a substantial yard is maintained, which makes it possible to Stockpile 2500 tons of oxidized ore, 1000 tons of hard sulphide ore, and 1000 tons of gougy sulphide ore. The bins are fed by a 60-hp. caterpillar bulldozer from the stockpiles. Fig. I shows the method and

Jan 1, 1943

By Fred Wise, C. W. Wark

The Getchell mine, a comparatively recent gold discovery, is in the old Potosi mining district, Humboldt County, Nevada. All ore is mined from open pits using Diesel shovels and gasoline trucks. Two types of ore are mined, oxide and sulphide. Oxide ores are siliceous or shaly; sulphide ores either gougy, siliceous or shaly. In mining, the sulphide ore is segregated into hard and gougy ore, because the gouge must be dried before it can be crushed. The oxide ore is not segregated. General Metallurgy The oxide ore is the leached outcrop that overlies the sulphide ore body. All of the iron and arsenic sulphides found in the sulphide ores have been thoroughly oxidized. The sulphur and arsenic have been leached out while the iron is present as hematite and limonite. The oxide ore is amenable to direct cyanidation, and although no visible free gold is present it is released at a comparatively coarse grind. The sulphide ore is more refractory, since the gold is all of micron size and some of it is locked in fine sulphides. It has been established microscopically that the free gold is coated with an impervious shell, which is probably some form of iron compound. The sulphide ores contain from 1.5 to 2 per cent of arsenic, which is present as orpiment and realgar with minor amounts of arsenopyrite. The orpiment and realgar are of a later mineralization than the gold and are barren. Roasting and washing of the sulphide ore prior to cyanidation is necessary to eliminate , the arsenic, to alter the gold coating from an insoluble form to a soluble form, and to release the gold associated with the fine sulphides. Washing the ground calcine in weak sulphuric acid, trona, or hot water are beneficial in the order named, but to date the water is the best from an economic standpoint. Metallurgical tests as well as plant practice indicate that a certain amount of gold cannot be recovered even by going beyond economic practices. On this basis metallurgical results are viewed in the light of an optimum tailing rather than a percentage of extraction. Milling Practice Crushing (Fig. I).—There are three primary ore bins, one each for oxide ore, gougy sulphide ore, and hard sulphide ore. In front of each bin a substantial yard is maintained, which makes it possible to Stockpile 2500 tons of oxidized ore, 1000 tons of hard sulphide ore, and 1000 tons of gougy sulphide ore. The bins are fed by a 60-hp. caterpillar bulldozer from the stockpiles. Fig. I shows the method and

Jan 1, 1943

By W. A. Bruce

THIS paper presents the essentials of a mathematical method of studying the pressure behavior of an oil reservoir as the fluids are withdrawn. Methods are shown whereby the behavior of a reservoir can be used to predict the future relationship between withdrawals and pressure. The oil reservoir is considered as a small part of a large porous continuum, which contains for the most part water. The expansion of this water resulting from a decrease of pressure in the oil zone causes a movement of water that is an important factor in the relationship between fluid withdrawals and reservoir pressure. Examples are given to illustrate the methods of pressure prediction for a circular reservoir containing undersaturated crude, the pressure distribution over a field that has a uniform production rate, the effect of a fault line near a field, the mutual interference of near-by fields, and the perturbing influence of a gas cap. Applications of the analysis to sell-spacing problems are shown.

Jan 1, 1942

By A. D. McMaster, M. L. Gimpl, N. Fuschillo

Low-power, high-speed, radiation-resistant, monolithic thin-film integrated circuits require thin-film resistors of high sheet resistance which are compatible with the processing requirements for monolithic silicon devices. For this purpose the structure and electrical resistance as a function of temperature were investigated for Nichrome-silicon monoxide cermet films on passivated single-crystal silicon substrates. A suitable material was found to be a mixture of 50 vol pct Nichrome-50 vol pct silicon monoxide. This mixture, when evaporated onto 300°C substrates. showed a low temperature coefficient up to 400°C, at both 270 and 1100 ohms per square. Resistors deposited on room-temperature, 300°, and 500°C substrates were heat-treated at 600°C in air for short periods of H- seed, low-power, radiation-re sis tant, monolithic thin-film integrated circuits'-3 require thin-film resistors of high sheet resistance and low temperature coefficients of resistance which are compatible with the processing requirements for monolithic silicon active devices. The present investigation is directed towards the replacement of the Nichrome in the sandwich-type SO2-Nichrome-SiO resistors, currently used for DTL monolithic thin-film integrated circuits, with codeposited Nichrome-SiO cermet resistors. These Nichrome-SiO resistors are provided with an overcoat of SiO and are deposited onto thermally oxidized silicon wafers used in monolithic mi-crocircuit technology. The cermet resistors are shown to provide a lower temperature coefficient of time totaling up to 45 min. Electron nzicrogvaphs and diffraction patterns were taken of the as - deposited and heat-treated films. These results were then compared with electrical-resistance measurements for each condition. The room-temperature and 300°C substrate films showed a large decrease in resistance. The 500°C substrate film remained stable throughout the entire cycle. Electron-microscopy results show that the increased stability after heating to 600°C is largely due to crystallization of the films which begins in the vicinity of 400°C. The slight oxidation of the films when heated in air had a negligible effect on the resistance. The resistance of the film was a function of the substructure. resistance and a more versatile compatibility with silicon monolithic processing. Previous work3, on chromium-SiO coevaporated cermet resistors on glass and ceramic substrates without overcoats of SiO has been reported. For the present application, the thin films should possess the following ideal requirements: a) a temperature coefficient of resistance as near zero as possible up to 300°C; b) have a sheet resistance higher than 200 ohms per sq to conserve space on the chip; c) be easy to evaporate; d) be oxidiation- and radiation-resistant; e) show long-time stability at 100°C; } be stable in N2 atmospheres up to 600°C for short periods of time (5 to 45 min). Requirements a to e are well-known. The last requirement was optional and was due to a mask elimi-

Jan 1, 1967

By S. C. Miller, R. D. Carter, H. G. Riley

A merhod based on theoretical considerations is presented for estimating srabilized performance of gas wells from data obtained from short-term flow tests. Calculated results are presented for II gas wells evaluated with four-to five-hour flow tests. For seven wells on which comparison war available, satisfactory agreement between measured and calculated performance was obfained. Although the final results are favorable, diflerences in flow capacity k,h obtained from flow tests conducted at different rates were observed. This inconsistency did not seriously affect the calculated results in the test wells. INTRODUCTION In order to negotiate the sale of gas from newly drilled gas wells and to plan otherwise for the orderly exploitation of gas reserves, it is necessary to predict the long-term withdrawal rates physically possible for each well. Since low permeability gas wells do not stabilize except after a lengthy period of flow, the operator is unable, without wasting large amounts of gas, to measure directly the stabilized performance prior to contracting for sale of the gas and connecting the well to a pipeline. This problem is not new, but is one which has recently received considerable attention because of the rising importance of natural gas in the economy and new emphasis on the development of gas reserves. Thus, a method is needed for estimating stabilized performance from information obtained from short-flow tests in which stabilized flow is not attained. Such short tests would be suitable for wells not connected to a pipeline since the tests require only small amounts of time, labor and production of gas. Other requirements of such a method are short calculation time and use of a model of the gas well and its reservoir drainage area which is as consistent as possible with present theoretical, experimental and field knowledge. The method presented herein was developed in an effort to meet these requirements. It has been named the "two-Row method". In addition to the section on the evelopment and description of the method, a description of field application of the method and a discussion of well test results are in- cluded. Finally, there is a section on an additional consideration prompted by some of the test results. The Appendix contains certain details of the derivation of the two-flow method and a sample calculation for Test Well 1. MATHEMATICAL DEVELOPMENT AND DESCRIPTION OF THE METHOD The method is based upon the following assumptions: (1) The system can be described as a finite cylindrical gas reservoir produced by a single cylindrical well located at the center of the reservoir. Flow is strictly radial. (2) Flow obeys Darcy's law except very near the well where a quadratic flow law is assumed. (3) The system is also characterized by a laminar flow skin factor. The use of a quadratic or turbulent flow term is more fully justified in Refs. 1 and 2. When such a system is opened at the wellbore to constant rate flow from an initial condition of uniform pressure distribution, the radius or drainage quickly passes through the non-Darcy flow and "skin effect" region imrnediately surrounding the well. After the drainage radius or "transient" has passed beyond this region and before it reaches the external boundary, drawdown can be expressed by'." After the drainage radius reaches the external boundary, stabilized flow exists and drawdown is given by'," Eqs. 1 and 2 form the basis for the method. The validity of these equations will now be examined in the light of some well test data previously available. It has been found by actual measurements on gas wells that the relationship between drawdown and flow rate, that is the performance curve, is closely approximated by' Eq. 3 is a straight line on log-log graph paper. Eq. 2 and, for any lixed value of time, Eq. 1 plot as a curve which is slightly concave upward on log-log graph

By H. E. Wilborg, G. V. Henderson

This chapter deals with those sands employed by foundries for the manufacture of cores and molds used in the casting of such common metals as steel, gray iron, ductile iron, aluminum-based alloys, and copper-based alloys. Most metal is cast in "green sand" molds in which the sand mixture, consisting primarily of silica, clay, and water, forms a plastic material. This plastic material is rammed around a pattern in a split box called a flask, then the box is separated, the pattern is withdrawn, and the metal is poured. These green sand molds are broken apart when the metal has solidified. The sand is then remixed and a new mold made for the next casting. To a lesser extent silica sand is mixed with resins or oils and, after forming, the mixture is hardened by either heat or chemical reaction. In any case, the foundry mold or core is used just once before being broken up and remixed for manufacturing another core or mold. By far the greatest consumption in the foundry is of those types of sand high in silica, low in clay (under 0.5%), and occurring in numerous locations thoughout the United States. Consumption End Uses Foundry sands fall into four distinct categories: silica sand, lake sand, bank sand, and natural molding sand. Silica sand is a general term used to describe washed, graded, dried, and cooled clay-free sands which are found in many parts of the United States. They generally contain 90 to 100% silica but vary greatly in chemical analysis and grain structure. Lake sand is the sand dredged from Lake Michigan or dug from the dunes bordering that lake. Bank sand is composed primarily of silica grain but containing up to 5% clay. It is shipped in both the damp and dry state but the normal processing of bank sand consists only of drying and scalping. Occasionally two different grades of bank sand are blended prior to or during loading. Natural molding sand is a term used to describe a sand containing sufficient clay as mined to be suitable for molding purposes. Clay content may vary from 10 to 24% with the balance of the material consisting primarily of silica sand. These sands range in fineness from one peaked on the 20 mesh to one which contains more than 50% of the particles passing the 270 mesh. Lake sands are relatively coarse and are used in great quantities for the production of automotive castings. Bank sands and natural molding sands are gradually phasing out as the production of castings becomes more a science and less an art. The vast majority of foundry sands used today are those known as the silica sands and lake sands. Foundry sands are used in two basic ways. The majority are used in the green or uncured state with the balance used in the cured or set state. Green sand is grain bonded with clay and contains varying amounts of water to plasticize the clay in order to bond the grains together. Green sand is either naturally bonded or "synthetic" sand. The naturally bonded molding sands are generally found in the Hudson River Valley of New York State, Cumberland County in New Jersey, and in Tennessee and Mississippi. Synthetic sand is molding sand compounded in the individual foundry. It is made up primarily of graded silica sand, western or southern bentonite, water, and carbonaceous and/or cellulosic materials. Green sand is used primarily in making molds. The use of naturally bonded green sand has been declining for many years, primarily because naturally bonded sands are more difficult for the foundryman to find and control. Synthetic green sand, on the other hand, can be

Jan 1, 1983

By Fred D. DeVaney

DURING the past few years a radically new process for the magnetic roasting of iron ores has been investigated and developed by Pickands Mather & Co. and the Erie Mining Co. in the Erie laboratory at Hibbing, Minn. This process, originally devised by Dr. P. H. Royster of Washington, D. C., involves the use of a roasting technique quite different from older methods. It has now been demonstrated that iron-bearing materials can be roasted as effectively as by any previously known method, and at a much lower cost. The increasing shortage of highgrade iron ores in this country has accelerated the search for new methods that would permit low grade materials to be utilized. The concept of magnetically roasting low grade nonmagnetic ores such as the oxidized taconites and then separating such material magnetically has always had considerable appeal. The magnetic concentration idea is attractive because of the sharpness of the separations and cheapness of the method. Heretofore, however, the equipment and the processes available for the magnetizing-roasting -step have left much to be desired. The customary equipment available for reduction roasting has been: 1-multiple hearth furnaces, 2-rotary kilns, and 3-shaft type kilns. In addition, it is understood that some work has been done in magnetically roasting fine ores by a process using the FluoSolids principle, but little information on this process is available. The multiple hearth kiln has been used the most but first costs and operating costs have been high because of low capacity, high maintenance, and poor gas utilization. Magnetic roasting can be done in a rotary kiln, but the radiation losses are high and the conversion to magnetite is usually unsatisfactory because of poor contact between the gases and the solids. Of the shaft-type furnaces, probably the most efficient yet developed is that designed by E. W. Davis of the Minnesota Mines Experiment Station. This furnace was operated at Cooley, Minn., during 1934-1937 but was abandoned in 1937 because the operation was uneconomic. Heretofore the basic concept behind most magnetic roasting processes has been the idea of heating iron ore to a temperature of 800° to 1100 °F in a strong reducing atmosphere, preferably either carbon monoxide or hydrogen. Temperatures under 800°F were undesirable since excessive roasting time was required. Temperatures over 1100°F were avoided because of the danger of converting part of the iron to ferrous oxide which is nonmagnetic. In the new roasting process, the operation is carried on in a shaft furnace using a controlled atmosphere containing a low percentage of reducing gas. The temperature in the roasting zone is considerably higher than with the usual reducing gas and this speeds up the reduction time. Portions of the spent furnace gases are cooled and recirculated and this together with the good contact between ore and gas makes for high reducing gas utilization. High heat economy is secured by recuperating heat from the roasted ore by passing the cold reducing gases countercurrent to flow of ore. The heat transfer principle is similar to that employed in a pebble stove and to that used in the Erie Mining Co. furnace at Aurora, Minn., for pelletizing fine magnetite concentrates derived from taconite. The theory of controlled atmosphere during the roasting operation can best be appreciated by inspecting the equilibrium diagram of the Fe-C-O system shown in Fig. 1. An inspection of this diagram shows that in certain areas magnetite, Fe3O4, is the only stable form of iron. A further inspection of this table shows that if the proper ratio is maintained between carbon dioxide to carbon monoxide, such a gas will be reducing with respect to hematite, Fe2O3, and will be oxidizing with respect to both ferrous oxide, FeO, and iron, Fe. It should be kept in mind that the formation of ferrous oxide in a roasting operation is harmful, since this oxide is nonmagnetic; if it forms in any quantity, it will cause substantial loss of iron in the ensuing magnetic separation step. If a ratio of approximately three parts carbon dioxide to one of carbon monoxide is maintained, the resulting operation can be carried on at a relatively high temperature without fear of over-reduction. Specifically, most of the tests in the Erie furnace have been made at a temperature of 1500° to 1600°F, with an entrant gas containing approximately 5 pct carbon monoxide and 15 pct carbon dioxide, with the remainder largely nitrogen. It should be remembered that the ratios of carbon monoxide to carbon dioxide shown in Fig. 1 hold even though the bulk of the gas is an inert gas such as nitrogen. It may surprise many to learn that a gas containing as low as 3 pct carbon monoxide, and 12 pct carbon dioxide with the remainder nitrogen is an extremely effective reducing gas in the 1000° to 1600°F temperature range. The reducing gas is not limited to carbon monoxide, and mixtures of hydrogen and carbon monoxide may be used effectively, provided that a similar ratio is maintained between the reducing gases and carbon dioxide and water vapor. For a more detailed explanation of the theory involved, the reader is referred to U. S. patents 2,528,552 and 2,528,553. From a safety standpoint, the weak reducing gas used in the furnace offers an advantage. Its composition is such that it is well below the limits of explosion should air enter a hot furnace. This condition is not true with the usual reducing furnace, in which a gas rich in carbon monoxide or hydrogen is used. The general furnace design and method of operation may best be understood by an inspection of

Jan 1, 1952

By John V. Beall

Given: One large, but dated open-pit copper mine and processing plants, down about one third in production capacity in decade because of declining ore grade. Solve for: Restoration of former peak capacity without further losses during period of reconstruction. That is the problem that Kennecott's Utah Cop- per Division began solving in 1963. Most of the work went to a task force of engineers, several of whom were plucked out of line jobs and assigned to the so-called "Expansion Project" for the duration. There was no doubt from the beginning that effective cooperation between Expansion Project engineers and operators was critical to success, and personnel was chosen accordingly. This foresight paid off. The Expansion Project is all but complete, and the Utah Copper Division is returning to its rated capacity of 300,000 tons per year of copper metal.

Jan 6, 1967

Henry Carlisle: This is June, 1961, and my friend Arthur Bunker is on the other side of the tape recorder. I, Henry Carlisle, am pursuing the hobby of chronicling the interesting parts of the mining careers and the lives of well-known mining engineers. Now, Arthur, supposing we skip, for the time being, where you were born and brought up, and start right in with where you went to school. Arthur Bunker: I graduated from Taft School and then went to Yale, Sheffield Scientific School, and took a course in electrical engineering. I was quite fascinated by higher mathematics and physics, and this offered more in mathematics and physics than any other one of the engineering courses at that time. The school was only a three-year course, and so I graduated from there in 1916.

Jan 8, 1963

By S. Yukawa, M. J. Sinott

The low solubility of bismuth in copper and its segregation at copper grain boundaries with resulting embrittlement is well known.1-3 The heterogeneous diffusion of liquid bismuth into polycrystalline copper has also been studied.4 This work presents some further observations on these phenomena using high-resolution auto radiography to detect the presence of bismuth in copper polycrystalline and bicrystal specimens. The auto radiographs were made by the stripping film technique which produces an autoradiograph that is intact with specimen surface and can be examined simultaneously with the under-lying microstructure. The details of the technique have been described previously.3 In one series of experiments, high-purity copper was alloyed with 0.02 to 0.03 wt pct radioactive bismuth obtained from the Oak Ridge National Laboratory. Melting and casting was done under a

Jan 1, 1960

By David J. Selleck, Eugene P. Pfleider

9.61. Introduction. Open pit mining methods produce more than 80% of all raw materials today in the United States. Much of this comes either from properties that formerly employed underground methods and were converted to open pit or from mines that ultimately must go underground when the economic stripping limit is reached. Generally the ores are transported over large vertical distances and techniques have been developed to do this economically. One approach is to combine conventional surface trucking, over relatively level roads, with the use of gravity and underground facilities. There are various modifications, depending upon the character of the orebody and the topography. Several schemes, as used in Sweden, Canada, and the United States, merit describing. One of the critical problems involved in the transference of ores from surface to underground is that of flow of bulk solids. Must it be crushed? Will it hang up in the transfer? What size and inclination of ore pass are required. How is the ore best fed onto the transporting device underground? As an example, the planning engineers for the Carol Project of the Iron Ore Company of Canada faced these problems in making their designs for the production and transportation of some 55,000 tpd from their Smallwood Mine in Labrador. Other operations, facing somewhat similar conditions, used other approaches of interest. This chapter describes some of the factors involved in the design, construction, and operation of a combination surface underground transportation system as applied to open pit mining operations. 9.62. Tunnel-Ore Pass Systems. The type, shape, dip, and size of the orebody, as well as the surface topographic features, generally estab-

Jan 1, 1968

By Frank Doxey

AT Dominion Steel & Coal Corp. it has long been recognized that continued mechanization of mine operations is necessary in the Pictou, Cumberland, and Sidney coal fields of Nova Scotia. The varied physical conditions in these fields call for special consideration of individual cases before planning is finalized. Because standard equipment cannot be procured which would operate successfully, many experiments have been necessary over the years to keep pace with the progress made in other countries. There are two mines, producing 2000 tpd, located in the Pictou coal field. The field is badly distorted and crossed by many faults. Seams are highly inclined and irregular and vary in thickness from 5 to 40 ft. Entries are difficult to maintain because of squeezing of the coal ribs and movement of the roof and pavement.

Jan 7, 1954

By F. D. Rosi

The slip and twinning behavior in extended titanium crystals were studied in some detail. The formation and appearance of coarse kink bands are discussed. Their crystallographic geometry was determined by X-ray analysis. A phenomenological interpretation of the complexities in kink band development is also presented. The critical resolved shear stress, coefficient of shear hardening, and plane of fracture were determined for several crystals extended at room temperature. THE slip and twinning elements observed in the room-temperature deformation of titanium were enumerated in a previous paper1 in which considerations were advanced regarding the nature and selection of these elements and their effect on the known mechanical properties of this metal. The present study concerns the crystallographic, microscopic, and mechanical aspects of flow in relation to the slip and twinning elements, and includes a prediction of slip systems, nature of slip and twin markings, inhomogeneities of plastic flow, and stress-strain characteristics. Unless otherwise noted, arc-melted titanium sponge (99.77 pct) was used in these experiments. The method of production of crystals, their dimensions, and the surface preparation for micrographic examination have been reported.&apos; For obtaining stress-strain characteristics, only those crystals which traversed the entire width of the specimen and were at least 8 mm in length were used. Tensile deformation of the crystals was performed with conventional grips for sheet specimens and a constant-stress loading beam, designed after the method of Andrade and Chalmers.&apos; The specimens were loaded by allowing sand to flow from a reservoir into a bucket suspended from the longer end of a balanced 6:1 lever arm at a rate controlled to load the specimen approximately 2 kg per min. Strain measurements were made using the Baldwin SR-4, bonded, resistance-wire strain gage, Type A-8, which permitted a reading accuracy of 2 microinches per inch. The formulas used in the evaluation of shear stress and shear strain, as in deriving the coefficient of shear hardening, are given by Schmid and Boasv n terms of the original orientation and change in length of the crystal. For calculating the critical resolved shear stress, the standard equation was used. The crystallographic nature of the unpredictable slip observed in a number of specimens was determined by the single-surface X-ray method of analysis as described in ref. 1. Experimental Results Prediction of Slip System: It was reported&apos; that room temperature slip in titanium takes place predominantly on {10i0} and in a <1120> direction, giving three potential slip systems. Hence, it should be possible to predict the operative primary system in the manner used by Taylor and Elam&apos; for alu- minum (i.e., the one with the greatest component of shear stress in the direction of slip). The stereo-graphic construction in Fig. 1 shows that this is true. In all cases, slip was found to occur initially on the (0110) plane and, in a number of cases, in the [21f0] direction. (The direction of slip was not determined for all orientations.) It follows that duplex slip can be expected when two slip systems are geometrically equally favorable for slip, as demonstrated in Fig. 2. In both crystals slip took place simultaneously on the (1700) and (olio) prismatic planes, as indicated by Fig. 2a and b. In the case of crystal B the movement of the specimen axis with increasing extension was toward the (10i0) direction, which represents the stable end orientation resulting from alternate slip on the [2ii0] and [1120] directions. This is analogous to the duplex slip process in face-centered cubic metals." Unpredictable Slip: In a number of crystals whose orientations were well within the operation of a single slip system, secondary slip occurred on planes not predicted by the criterion of maximum shear stress. Examples are shown in Fig. 3. In addition to

Jan 1, 1955

By J. M. Vonfeld, V. D. Hanson, W. K. Heinlein

TWO overfeed drum-type separators using a suspension of magnetite in water as the separating medium have been installed in the Champion No. 1 preparation plant of the Pittsburgh Coal Co., Division of the Pittsburgh Consolidation Coal Company. A simplified flow diagram of the Champion No. 1 coarse coal circuit is shown in Fig. 1. A railroad car rotary dump, capable of handling 20 cars per hr, discharges the coal into a hopper. Coal is fed from the hopper of the rotary dump to a 60-in. belt conveyor by two reciprocating pan-type feeders. The belt conveyor feeds the run of mine coal onto the main raw coal screen. The first deck is fitted with 3 in. round hole perforated plate which removes the 3x0-in. coal. The second deck of the screen is equipped with 6-in. lip screens which separate the remaining coal into +6 in. and 6x3 in. The +6 in. coal is transferred to a shaking picking table and then discharged onto the loading boom. The 6x3-in. coal is fed to the No. 2 dense medium vessel for cleaning. The clean coal from this unit flows to a vibrating screen where the medium is removed and the coal sized into 6x4 in. and 4x3 in. The 6x4-in. coal is transferred to the loading booms and the 4x3-in. is conveyed to the clean coal classifying screens to remove degradation. The 3x0-in. raw coal is blended in a system composed of three storage bins having a total capacity of 400 tons. From these bins the coal is fed onto a 48-in. flight conveyor by three reciprocating feeders. This flight conveyor delivers the coal to the Rheo plant. The 3x0-in. coal is then cleaned in two 48-in. Rheolaveur launders each equipped with two Rheo boxes. The first set of boxes removes a primary refuse and the second set produces a middling product which is recirculated in the launders. The cleaned coal and water overflows from the primary launders to two sizing and dewatering shaker screens where the water and - 3/8-in. coal are removed for further cleaning in the Rheo fine coal plant. The 3x3/8-in. clean coal is combined with the 4x3-in. coal from the No. 2 vessel, sized into 4x2 in., 2x1 1/4 in., and 11/4x3/8 in., then sent to the loading booms. The 3x0-in. refuse from the primary launders is normally fed to the No. 1 dense medium circuit, but any portion or all of this material can, as an alternate, be processed in the Rheo rewash launder. The primary refuse when fed to the dense medium circuit is screened into 3/8x0 which is cleaned in the fine coal plant and 3x3/8 for cleaning in the No. 1 vessel. The clean coal from this vessel is passed over a vibrating drainage and rinse screen, then sluiced to the rewash sizing shaker where it is dewatered and sized for loading. The installation of a pilot plant separating vessel using a suspension of magnetite in water was pro-

Jan 1, 1952

By J. W. Rutter, K. T. Aust

Competitive growth of recrystallized grains into striated single crystals of zone-refined lead produced preferred orientations of the coincidence type after annealing at 175°C, but not at 300°C. This preferred orientation development was found to result from the higher mobility and lower energy of large -angle coincidence boundaries compared to large-angle noncoincidence boundaries. These mobility and energy difference provide a useful basis for explaining many cases of preferred orientation observed in annealed metals. DURING a study1 of the mobility of grain boundaries in zone refined lead, it was observed that the temperature dependence of the grain boundary migration rate is smaller for large-angle, coincidence-type grain boundaries than for large-angle noncoincidence ("random") boundaries. This observation is illustrated schematically in Fig. 1, which shows log rate of grain boundary migration vs reciprocal absolute temperature for two large-angle boundaries. The upper curve refers to a coincidence-type boundary of the sort discussed by Kronberg and Wilson.2 While the mobilities of these two boundaries are similar at 300°C, the coincidence-type boundary has a greater mobility, by a factor of about: four, at 175°C than does the noncoincidence boundary illustrated in the lower curve. The observation described above suggested a simple experiment for testing the role of grain boundary mobility in the development of preferred orientation in annealed, high-purity lead. On the basis of this observation, it would be expected that if a large number of grains of different orientations are growing competitively, and if relative grain boundary mobility is a controlling factor, then the specimen procluced by boundary migration will contain crystals having random orientations if annealing is done at 300°C, while preferred orientations of

Jan 1, 1962

By D. N. Sutherland

The cassiterite flotation circuit at the Renison concentrator has operated with increasing levels of automatic control since late 1981. The circuit operates at high levels of recycle which make it difficult to control manually so that the initial control strategy was to stabilize the mass flows around the circuit prior to circuit optimization. Control of the air supplies to the flotation banks was introduced to achieve this stabilization and initial results indicated that some success has been achieved, as indicated by improvements in the overall circuit performance and reduction in the stream variabilities. These improvements were gained through control of the major recycle to the roughing stage which dominates the circuit behaviour. Feed-forward control of collector addition has also been introduced successfully, but attempts to control the final concentrate grade have been less satisfactory. This work is continuing and the objective of current activity is to include the three cleaning stages in an overall circuit control scheme as a first phase of metallurgical optimization of the operation.

Jan 1, 1984

By T. C. Atchison, W. I. Duvall, D. E. Fogelson

Dr. Clark has given an excellent resum6 of the recent theoretical re- search work that has been done on the generation and propagation of stress waves in various types of media.1 Unfortunately the dynamic properties of rock do not fit the assumed properties of the idealized media. As a result, none of the theoretical solutions for the generation of stress waves by applied pressures to interior boundaries of the spherical or cylindrical cavity appear to be completely satisfactory to explain experimental data. This fact does not mean that the theoretical work is in error. It simply means that the appropriate dynamic properties of rock have not been taken into account in the present theoretical derivations. Because of the difficulty of theoretically accounting for the crushing and cracking that takes place when explosives are detonated in spherical or cylindrical cavities, it seems advisable that future theoretical studies should attempt to separate the generation of stress waves from the propagation of stress waves. Perhaps if these two problems are separated, we may be able to solve one of them satisfactorily so that theory and experimental data agree. Dr. Clark&apos;s resumé of the research work on scaling and crater formation by point charges is aimed at obtaining a basic understanding of the physical theory of rock breakage, wave generation, and wave propagation. A cursory reading of Dr. Clark&apos;s paper shows that there is very little information summarized that can be used by the practicing engineer in the field of blasting. When problems become too difficult to solve theoretically it is always helpful to take the empirical approach and collect experi-

Jan 1, 1967

By Jed H. Mosgrove, Paul M. Budzak

INTRODUCTION Good mine ventilation starts at the drawing board. A coal mine will either be adequately ventilated or not, depending upon how good a job was done in the planning. Poor planning will not only result in inadequate ventilation, but may also lead to low productivity. A mine foreman cannot achieve high productivity while trying to maintain 4.2 m3/s (9000 cfm) of air at his last open crosscut. PLANNING What will provide adequate ventilation? A rule of thumb in laying out a coal mine is to provide enough entries in the main entry system to keep air velocities at 2.5 m/s (500 fpm) or less; at this rate of flow, good ventilation may be expected. The intake and return air courses of a coal mine do not improve with age. Entries will deteriorate, ribs will slough, roof will fall, and clutter will accumulate. The resistance offered to air travel will continue to increase with each passing year. Providing enough airways for the present and the future cannot be emphasized too strongly. In order to better appreciate good planning and bring it into better focus, assume a hypothetical mine has the following features: entries 1.5 m (5 ft) high and 6.1 m (20 ft) wide; one intake and one return entry; the mine development is 4.8 km (3 miles) from the intake and upcast ventilation shafts; and the ventilation requirement is 141.6 m3/s (300,000 cfm) . Air travel of 4.8 km (3 miles) into a mine and the same distance back to the fan is not unusual. However, to do so with one intake and one

Jan 1, 1981

By AIME AIME

THE American copper mining industry is threatened with disintegration and destruction. This threat is not one which may only materialize in the distant future. The destruction has already commenced. At present, many domestic mines have already closed down. .Others are operating on a greatly curtailed basis. Those operating are doing so only at tremendous loss-because the selling price of copper is below the cost of production and rapidly diminishing financial resources and credit clearly forecast the early suspension of operations in most of the industry. Unemployment in the mining districts is increasing by leaps and bounds. While this condition is due in some measure to the present unsatisfactory state of business in general, the copper industry, unlike other lines of business, sees no possibility of recovery without definite aid. The fundamental, underlying cause of present conditions in the American copper mining industry is the unprecedented rise in foreign production, and the basic change which has taken place in the world copper industry in the last three years. The American industry had begun to feel this change early in 1929 long before the general collapse of business in the fall of that year. While the distress in the industry is undoubtedly aggravated by the present economic depression, it nevertheless would have arrived and. have been present at this time, regardless of general business conditions.

Jan 1, 1932