Search Documents

By Jay R. Burns

TWO magnesium sand-casting alloys are commonly favored in the United States. These are referred to as H and C alloys (Dow Chemical Co.) or 4Mz65 and AM260 alloys (American Magnesium Corporation). Both are of an aluminum-zinc-manganese type. H alloy contains 6 per cent aluminum and 3 per cent zinc and is characterized by relatively good toughness. C alloy contains 9 per cent aluminum and 2 per cent zinc and is employed where high yield strength and pressure tightness are required. Examination of magnesium castings poured in Europe and England,'J however, demonstrates that H and C alloys are replaced by .%8 and AZ~I, respectively. That is, A8 reprcsents a magnesiurn alloy having superior toughness while A291 is suitable for high yield strength and pressure-tight applications. The basic difference between the .Ameri-can and European alloys is that the foreign alloys employ a somewhat higher aluminum content and less zinc than is found in H and C alloys. The mechanical properties of the alloys, when divided into high and low-aluminum groups, are quite similar, as determined from separately cast foundry-control type test bars ,In this paper the decrease of mechanical properties caused by any given amount of microporosity is measured for H, C, A8, and AZ~I alloys. The relative susceptibility of each of the four alloys to microporosity is also determined. Typical mechanical properties and nominal chemical analyses are reported in Table I. The apparent similarities between the two alloys in each group have led to considerable discussion as to the relative merits of each of the compositions.2 Popular points for comparison between the alloys have been the relative susceptibility of each of the four alloys to microporosity6 and the effect of this microporosity on their mechanical properties. It is primarily intended, in this paper, to show the

Jan 1, 1946

By Walter C. Lawson

THE first plans were made in 1930 for the mining by open-pit methods of the low-grade disseminated ore body now known as the Morenci open pit. It was not until 1937, however, that final plans were completed and the decision reached to go ahead with the actual development. Many advances were made in open-pit mining equipment during that period of seven years, therefore the ways and means of attack actually used were very different from those originally contemplated. Diesel-operated bulldozers, big trucks, full-revolving electric shovels and electric churn drills constitute the major equipment finally purchased for this modern stripping operation. Neither the bulldozers nor the big trucks were considered in the early operating plans, for at that time bulldozers had not been used in mining operations and the trucks as yet were undesigned. The full-revolving shovel and the churn drills were not new developments in open-pit equipment, but both had undergone remarkable manufacturing transition, emerging far more efficient and far more dependable than their predecessors. Even the ordinary drilling steel had been fundamentally changed. It was now threaded so that the dulled cutting bits could be detached and replaced, thus eliminating the costly transportation of long bars to the drill shop for resharpening. The decision to use trucks in the preliminary stripping operations was based on the desirability of this type of transportation to economically prepare the pit for rail haulage, coupled with the success that the contractors had had with them on other big excavation jobs. There are at present no plans for the use of trucks beyond the preliminary work, inasmuch as rail haulage will give lower costs after roadbeds and benches are once established. For many years to come, both ore and waste haulage will be over favorable grades. The average length of ore haul between the pit and an assembly yard will be about two miles. From there the crushers will be 1 1/3 miles with a connecting track grade of 0.4 per cent. The upper ore benches of the pit will be reached from the yard by a series of four switchbacks over a ruling grade of 4.0 per cent compensated for curvature. Tracks to the dumps will be on 0.2 per cent grades.

Jan 1, 1938

By D. L. Creighton, S. A. Hoenig

The field-emission microscope has been adapted for the study of microcrack growth during the early stages of fracture in metal wires. Cracks as small as 6 1 in length can be detected and their growth can be followed to specimen failure. The system is quite useful in searching for microcracks since only sharp-edged surface defects will emit electrons under the experimental conditions. THE conditions leading to brittle fracture were discussed a number of years ago by Griffith1 and the term Griffith Cracks is often used for the small surface cracks which are responsible for brittle fracture. Griffith's theory has been modified by stroh2 and more recent results on metals are discussed by Allen,3 pp. 123-40. At present the phenomenon is not completely understood but there is general agreement that at least in certain materials the sequence leading to brittle fracture involves several stages. The initial microcracks are present because of cooling or working stresses, Hahn et al.,3 p. 95. When a stress is applied to the specimen the cracks grow slowly until the release of stored elastic energy is large enough to accelerate the crack and provide the necessary surface energy for crack growth. At this point the growth rate appears to increase rapidly to some new equilibrium velocity, and failure occurs. Since the microcracks are usually about the size of a single metallic grain (Ref. 3, p. 99) it is not easy to find them and it is very difficult to follow their growth under stress. This paper will report on the use of a cylindrical field-emission microscope for observation of the formation and growth of microcracks. I) THE FIELD-EMISSION MICROSCOPE The field-emission microscope (FEM) has a high magnification and resolution and is almost uniquely suited for observations of microcracks. Since the FEM is relatively new as a metallurgical instrument, a short description will be given here. Normally metals at room temperature do not emit electrons; however in the presence of a strong electric-field gradient, electrons can tunnel out through the reduced potential barrier. Since this tunneling is a function of the local field gradient and the local work function, the emitted electrons can be used to produce a highly magnified image of the surface by allowing them to strike a phosphor screen. Because the electron emission is dependent upon the local field gradient, smooth surfaces emit few electrons except at very high fields. On the other hand cracks, extrusions, or other surface defects, having sharp edges, emit strongly since the field gradient is very high in the vicinity of these defects. This indicates that the FEM should be most useful for detection of microcracks on otherwise smooth surfaces. A field-emission microscope was first used by Muller4 in 1936 for observation of metal surfaces, and recent reviews have been given by Muller5 and Gomer.6 The instrument has been used for metallurgical studies in the area of surface diffusion,= recrystallization,7 and grain growth 8 (Ref. 8 is directed specifically at metallurgists). In the work of Muller4,5 and Gomer 6 the specimen was in the form of a sharp metal point at the center of a phosphor-coated glais sphere. The impact of the emitted electrons on the phosphor produced a highly magnified image of the specimens. Such a system is not practical for applying a controlled stress to the specimen and a cylindrical geometry has been used in this investigation. This allowed the application of a controlled tensile stress to the wire specimen. Normally a cylindrical FEM geometry produces magnification only in the radial direction. This is the case because a smooth wire at the center of a cylinder produces a purely radial electrical field. However, if there is a break in the smooth surface of the inner cylinder, the field near the break becomes three-dimensional and the area of the break is highly magnified. The reason for this is clear if it is recalled that the field gradient depends on the relative radii of the inner and outer cylinders; if a crack forms, its edge radii are of atomic dimensions and a very high field gradient is formed near these crack edges. Since the electrons receive most of their acceleration near the crack edge and are always traveling perpendicular to the field lines, they tend to spread out and produce the magnified image observed in the cylindrical field-emission microscope. 11) BRITTLE-FRACTURE STUDIES A) Experimental Apparatus. The geometrical arrangement chosen was that used earlier by Gifford

Jan 1, 1965

By A. P. Svenningsen

IN the early stages of design it was not considered necessary that separate crushing plants be built for the new sulphide concentrator and smelter until sometime in the future. The plan was to use the existing crushing facilities for both oxide and sulphide ore. A few additions were contemplated for the existing plants, such as increased bin capacity, and possibly two new secondary crushing units. The more the problem was studied and discussed with the plant operators, the more it became evident that it was complex. It involved the classification of different kinds of ore from the open pit mine -sulphide, oxide and mixed-and how best to separate them so that each kind of ore was given the proper processing and treatment. It also involved the problem of keeping the different ores from being contaminated in bins, hoppers and chutes. Added to these, transportation became complicated and would involve additional handling and loading of ore from crushing plants to conveyors, to bins, and finally to railroad cars which were to be hauled to the concentrator and dumped into the fine ore bin. General In the early part of 1951 it was decided that the concentrator be constructed with ten grinding units instead of seven as originally authorized. The smelter was to be increased proportionally and naturally also the overall tonnages of ore to be handled by the new sulphide plant. Due to this increase in plant capacity and the larger tonnages involved, the difficulties which would arise by using the existing crushing plants were increased to a point where it became evident that the building of new crushing plants for sulphide ore exclusively was technically, as well as economically, advantageous. Authorization was, therefore, given by the company to construct new crushing plants to handle 30,000 tons of ore per day, and capable of reducing the run-of-open-pit ore to the proper size feed for the 10x14-ft rod mills in the concentrator. The ore, mined in the open pit, sometimes comes in pieces as large as 6 to 7 ft diam. The rod mills may call for ore crushed to 3/4 in. The large .size of ore delivered from the open pit determined that a 60-in. gyratory crusher be used as primary breaker. Such a crusher will have a capacity considerably in excess of 30,000 tons per day. The crusher will be a single discharge unit driven by a 500-hp electric motor through a tear coupling and a floating shaft. This type of drive has proven successful at a number of other crusher installations which our company has operating in the United States, Mexico and South America. The tear coupling will protect both the crusher and motor against damage in case of overload. No new features are incorporated in the design of the crusher itself, except that the, discharge chute is made the full width of the crusher with parallel sides instead of the usual converging sides. This change in detail should eliminate, a feature which has been a bottleneck in some of the operating plants and has caused loss of production due to ore hanging up and blocking the chute. The secondary crushing plants will have three 7-ft standard Symons cone crushers and six 7-ft short head Symons crushers. Between the primary and secondary crushing plants a coarse ore bin will be constructed with a nominal draw-off capacity of 30,000 tons of ore. The standard Symons and the short head Symons will be in separate buildings. All the crushing plants and the coarse ore bin are interconnected with conveyor belts for transporting the ore to the crushers at the tonnage rate desired. The final product of the new crushing plants is produced by the short head crushers. It will be delivered onto a conveyor belt leading to the top of the fines ore bin in the concentrator. A separate conveyor belt running the full length of the fines ore bin and provided with a movable tripper of rugged design will discharge the sulphide ore into the bin. The concentrator bin is planned and designed so that the installation of this additional conveyor will not interfere with the operation of the two railroad tracks on which crushed ore is brought from the existing oxide plant. Thus when completed the bin can be filled simultaneously by ore from the new crushing plant and by ore from the existing leaching plant.

Jan 1, 1952

By John E. Conley

THE production of lime by the burning or calcination of limestone, including all varieties from true dolomites and magnesian limestones to high-calcium types, continues as one of the essential basic industries. In 1937, a total of 4,124,165 tons of lime valued at approximately $30,091,000 was produced in the United States.1 The importance of any contribution to our knowledge on the mechanism of the lime-burning process is, therefore, obvious. The burning of limestone has been practiced since the earliest days of our civilization, and the literature on the various phases of the industry is voluminous. Calcination of the various types of limestone has been studied by many previous investigators, but some points require additional research. Each occurrence of limestone provides an opportunity for specific investigation, because the characteristics of stones in different localities vary Knibbs13 has given an excellent discussion of available data on the theory of lime and magnesia burning. The thermal dissociation of calcium carbonate has been studied more recently by Johnston,2 Smyth and Adams,3 Ralston et al.,4 Furnas,5 Potapenko,6 and Southard and Royster.7 As a result of these investigations the dissociation temperature at atmospheric pressure has been established as 894º to 898º C. The dissociation of dolomite and magnesite, although studied by several investigator,4, 6, 8 still requires clarification by additional data. The calcination of limestone was studied for a twofold purpose: (1) to extend our information on the mechanism of the lime-burning process and (2) to standardize the conditions for calcination of high-calcium and dolomitic limestones for subsequent hydration tests on the lime.

Jan 1, 1939

By Emerson F. Heald

A systematic study was made of new and old data on chemical activities in Fe-Pt alloys at elevated ternperatuves. Experimental results may be expressed in terms of the excess free energy using Least-squares analysis of the data gave the following values for the constants: for the temperature range 1130° to 1350°C, and tentatively to 155O°C, B = -3.326564 and C = 0.221051; for the temperature range 650" to 850°C, B = -2.555690, C = 1.762735, and D =0.097196. In the study of iron-containing silicate systems, it is sometimes desirable to have a direct experimental measure of the activity of iron in the system. The well-known solubility of iron in platinum, often a headache in experimental work on iron compounds under reducing conditions, can be used to advantage in this respect. If the activity of iron in an Fe-Pt alloy in equilibrium with the silicate is known as a function of the composition of the alloy, chemical analysis of the alloy will give a knowledge of the activity of iron in all of the phases in the system. The present study was undertaken in order to elucidate the characteristics of Fe-Pt alloys as iron activity indicators. This work is intended to tie together some previous work, which may be summarized as follows. Larson and Chipman1 determined the activity of iron in Fe-Pt alloys at 1550°C by equilibrating platinum metal with calcium oxide-iron oxide-silica melts of known iron activity. Compositions of the resulting alloys were determined by chemical analysis. A similar study was carried out by Taylor and ~uan,' who worked at 1300°C. They brought the Fe-Pt alloys into equilibrium with iron oxide under conditions of known partial pressure of oxygen, and thus, from the work of Darken and ~urr~,~ conditions of known iron activity. Compositions were determined indirectly, by following the change in weight of the sample. Sundaresen et el* used the electromotive force of cells in which the alloy formed one electrode in order to measure the activity of iron in the alloy at 650" and 850°C. These temperatures were chosen to be above and below the first-order phase transition which takes place upon the ordering of Fe3Pt and the second-order transition which occurs upon the ordering of FePt3. EXPERIMENTAL 1) High Temperatures. The starting materials used were thin platinum foil, about 0.002 mm thick, and Fisher Certified reagent ferric oxide, Fez03, which had been heated for 24 hr at 1000°C. An intimate mixture of 80-mesh Fez03 and platinum platelets was placed in a thin platinum foil envelope. The latter was suspended from thin platinum wires in the hot zone of a vertical-tube, platinum-wound furnace of the type described by Muan and ~sborn.~ A capillary gas mixer similar to that used by Darken and Gurry3 was used to prepare a precisely known mixture of carbon dioxide and hydrogen, which was allowed to flow upward through the furnace tube. The partial pressure of oxygen in contact with the sample was thereby fixed at a value which was calculated from the charts prepared by porter? Temperatures were measured with a Pt-10 pct Rh-in-platinum thermocouple, which was calibrated using the melting points of gold (1062 .@C) and diopside, CaMgSizOB (1391.5"C). Temperature control was maintained to within i3"C with a Geophysical Laboratory proportional controller, using the furnace resistance as the sensing element. Samples were quenched by passing a small current through the platinum suspension wires, allowing the sample to drop into a bath of dibutyl phthalate at the bottom of the furnace tube. Prior to chemical analysis the samples were washed with acetone and dried. 2) Chemical Analysis. It proved possible, in almost all cases, to separate the Pt-Fe platelets physically from particles of iron oxide. The platelets were dissolved in a small volume of aqua regia, evaporated to dryness, and redissolved to 0.1 M HC1. In order to determine iron in the platinum alloy potentiometrically, it is necessary first to remove the platinum. A 10-cm column of Amberlite IR-120 cation exchange resin in the hydrogen form provided separation quickly and quantitatively: The mixture of iron and platinum in 0.1 M HC1 was added to the top of the column, and washed with about 100 ml of 0.1 M HC1. Under these conditions, the iron, principally in the form of cations such as FeC1" and FeCl;, is held quantitatively in the uppermost centimeter of the column. The platinum, in the form of anions such as PtC&- , is washed through without being adsorbed. After a qualitative test with stannous chloride indicated all of the platinum was removed, the iron was

Jan 1, 1968

By C. B. Alcock, L. I. Cheng

By the use of radiochemical methods for the study of the gas-liquid equilibria at low temperature, and for the determination of the sulfur contents of metal beads which had been equilibrated with H2S/H2 mixtures of known sulfur potential, it has been possible to obtain the liquid solubility and the free energy of solution of sulfur in liquid tin and lead at temperatures between 500°and 680°C. THE gas-liquid equilibrium method has proved in the past to be most successful in the determination of the thermodynamic behavior of dilute solutions of sulfur in liquid metals.1,2 One of the basic requirements for success with this method is that the volatility of both the metal and its lowest sulfide should be small, otherwise sulfide will be deposited at the cool end of the furnace, where it may react with the outgoing gases to form either sulfur-rich lowest sulfide or higher sulfides. The resultant value of the apparent equilibrium constant will then be lower than the correct one. This argument applies even at sulfur potentials below that in equilibrium with a separate condensed phase of the lowest sulfide at the reaction temperature, T. The mass of sulfide which is deposited at the cold end of the furnace, and hence the extent to which further reaction occurs with the outgoing gases, depends on the time taken for equilibrium to be reached between metal and gas. Since this will depend principally on the bulk of the metal phase which is used, one should clearly attempt to uie as small metal samples as possible. These considerations are important in the study of dilute solutions of sulfur dissolved in liquid tin and lead which both have moderately high vapor pressures as metals and form volatile sulfides. The limit on the size of the metal samples which may be used is set chiefly by the difficulties of analysis for very small amounts of sulfur. The oxygen or carbon dioxide combustion method, followed by iodimetric determination of the sulfur dioxide which is formed,has been found to be successful for the determination of small amounts of sulfur in copper, iron, cobalt and nickel.4 This method was unsatisfactory for sulfur dissolved in tin and lead, mainly because the sulfur dioxide was to some extent absorbed by the copious tin or lead oxide deposits which were formed on the walls of the combustion tube. Furthermore some of the sulfur was found to segregate on the surface of the beads as flaky sulfide crystals which would easily be lost in the transfer of a bead from a boat in the gas equilibration apparatus to one in the combustion apparatus. Oxidation in aqueous media to sulfate ion followed by precipitation as barium sulfate was, therefore, adopted as the analytical procedure. The gas-metal equilibrium experiments were all carried out with radioactive sulfur and thus the analysis involved the counting of barium radiosulfate. Furthermore the use of the radioisotope meant that the approach to the gas-metal equilibrium could be followed continuously by gas counting.' The metal beads were held separately in glass crucibles during equilibration and were transferred from the furnace to the beaker for dissolution in nitric acid still in the crucibles, and thus the possibility of sulfur loss by detachment of the sulfide segregates was eliminated. The temperature range of this investigation was 500° to 680°C. EXPERIMENTAL APPARATUS AND METHOD The apparatus consisted of two furnaces placed in series in a gas recirculation system, Fig. 1. One furnace F1, which was vertical was used to heat the alumina crucible, A, holding six metal beads in separate glass crucibles. The beads weighed between 300 and 700 mg each. The crucible assembly was introduced and removed from the furnace mechanically under a stream of oxygen-free argon. The other furnace, F2, was horizontal and was used to heat a cobalt Co9S8 mixture, held in an alumina boat, and made with radiosulfur containing about 1/2 millicurie per g of sulphur. This mixture, which was finely powdered, was used as a source of known H2S/H2 mixtures6 for a given furnace temperature. The recirculation system also contained a gas re-circulation pump (P), an end window Geiger-Miiller counter (N)—placed downstream of F1 so as to monitor the H2S pressure in the gas leaving this furnace— a sample volume for chemical analysis of the gas phase (G), gas drying tubes (D), filling taps and other standard ancillary equipment. The gas sampling volume was principally used in the cali-

Jan 1, 1962

By John E. Conley

The production of lime by the burning or calcination of limestone, including all varieties from true dolomites and magnesian limestones to high-calcium types, continues as one of the essential basic industries. In 1937, a total of 4,124,165 tons of lime valued at approximately $30,091,000 was produced in the United States.' The importance of any contribution to our knowledge on the mechanism of the lime-burning process is, therefore, obvious. The burning of limestone has been practiced since the earliest days of our civilization, and the literature on the various phases of the industry is voluminous. Calcination of the various types of limestone has been studied by many previous investigators, but some points require additional research. Each occurrence of limestone provides an opportunity for specific investigation, because the characteristics of stones in different localities vary. Knibbs13 has given an excellent discussion of available data on the theory of lime and magnesia burning. The thermal dissociation of calcium carbonate has been studied more recently by Johnstonr2 Smyth and Adams,3 Ralston et al.,4 Furnas,6 Potapenko,6 and Southard and Royster.7 As a result of these investigations the dissociation temperature at atmospheric pressure has been established as 894" to 898' C. The dissociation of dolomite and magnesite, although studied by several investigators,4,6,8 still requires clarification by additional data. The calcination of limestone was studied for a twofold purpose: (I) to extend our information on the mechanism of the lime-burning process and (2) to standardize the conditions for calcination of high-calcium and dolomitic limestones for subsequent hydration tests on the lime.

Jan 1, 1942

By John E. Conley

The production of lime by the burning or calcination of limestone, including all varieties from true dolomites and magnesian limestones to high-calcium types, continues as one of the essential basic industries. In 1937, a total of 4,124,165 tons of lime valued at approximately $30,091,000 was produced in the United States.' The importance of any contribution to our knowledge on the mechanism of the lime-burning process is, therefore, obvious. The burning of limestone has been practiced since the earliest days of our civilization, and the literature on the various phases of the industry is voluminous. Calcination of the various types of limestone has been studied by many previous investigators, but some points require additional research. Each occurrence of limestone provides an opportunity for specific investigation, because the characteristics of stones in different localities vary. Knibbs13 has given an excellent discussion of available data on the theory of lime and magnesia burning. The thermal dissociation of calcium carbonate has been studied more recently by Johnstonr2 Smyth and Adams,3 Ralston et al.,4 Furnas,6 Potapenko,6 and Southard and Royster.7 As a result of these investigations the dissociation temperature at atmospheric pressure has been established as 894" to 898' C. The dissociation of dolomite and magnesite, although studied by several investigators,4,6,8 still requires clarification by additional data. The calcination of limestone was studied for a twofold purpose: (I) to extend our information on the mechanism of the lime-burning process and (2) to standardize the conditions for calcination of high-calcium and dolomitic limestones for subsequent hydration tests on the lime.

Jan 1, 1942

By John Henderson

FOR some time now the most commonly accepted description of liquid silicate structure has been the "discrete ion" theory, proposed originally by Bockris and owe.' This theory is that when certain metal oxides and silica are melted together, the continuous three dimensional silica lattice is broken down into large anionic groups, such as sheets, chains, and rings, to form a liquid containing these complex anions and simple cations. Each composition is characterized by "an equilibrium mixture of two or more of the discrete ions",' and increasing metal oxide content causes a decrease in ion size. The implication is, and this implication has received tacit approval from subsequent workers, that these anions are rigid structures and that once formed they are quite stable. The discrete ion theory has been found to fit the results of the great majority of structural studies, but in a few areas it is not entirely satisfactory. For example it does not explain clearly the effect of temperature on melt structure,3 nor does it allow for free oxygen ions over wide composition ranges, the occurrence of which has been postulated to explain sulfur4 and water5 solubility in liquid silicates. In lime-silica-alumina melts the discrete ion theory is even less satisfactory, and in particular the apparent difference in the mechanism of transport of calcium in electrical conduction8 and self-diffusion,' and the mechanism of the self-diffusion of oxygen8 are very difficult to explain on this basis. By looking at melt structure in a slightly different way, however, a model emerges that does not pose these problems. It has been suggested5" that at each composition in a liquid silicate, there is a distribution of anion sizes; thus the dominant anionic species might be Si3,O9 but as well as these anions the melt may contain say sis0:i anions. Decreasing silica content and increasing temperature are said9 to reduce the size of the dominant species. Taking this concept further, it is now suggested that these complexes are not the rigid, stable entities originally envisaged, but rather that they exist on a time-average basis. In this way large groups are continually decaying to smaller groups and small groups reforming to larger groups. The most complete transport data 8-10 available are for a melt containing 40 wt pct CaO, 40 wt pct SiO2, and 20 wt pct Al2O3. Recalculating this composition in terms of ion fractions and bearing in mind the relative sizes of the constituent ions, Table I, it seems reasonable to regard this liquid as almost close packed oxygens, containing the other ions interstitially, in which regions of local order exist. On this basis, all oxygen positions are equivalent and, since an oxygen is always adjacent to other oxygens, its diffusion occurs by successive small movements, in a cooperative manner, in accord with modern liquid theories." Silicon diffusion is much less favorable, firstly because there are fewer positions into which it can move and secondly, because it has the rather rigid restriction that it always tends to be co-ordinated with four oxygens. Silicon self-diffusion is therefore probably best regarded as being effected by the decay and reformation of anionic groups or, in other words, by the redistribution of regions of local order. Calcium self-diffusion should occur more readily than silicon, because its co-ordination requirements are not as stringent, but not as readily as oxygen, because there are fewer positions into which it can move. There is the further restriction that electrical neutrality must be maintained, hence calcium diffusion should be regarded as the process providing for electrical neutrality in the redistribution of regions of local order. That is, silicon and calcium self-diffusion occur, basically, by the same process. Aluminum self-diffusivity should be somewhere between calcium and silicon because, for reasons discussed elsewhere,' part of the aluminum is equivalent to calcium and part equivalent to silicon. Consider now self-diffusion as a rate process. The simplest equation is: D = Do exp (-E/RT) [I] This equation can be restated in much more explicit forms but neither the accuracy of the available data, nor the present state of knowledge of rate theory as applied to liquids justifies any degree of sophistication. Nevertheless the terms of Eq. [I] do have significance;12 Do is related, however loose this relationship may be, to the frequency with which reacting species are in favorable positions to diffuse, and E is an indication of the energy barrier that must be overcome to allow diffusion to proceed. For the 40 wt pct CaO, 40 wt pct SiO2, 20 wt pct Al2O3, melt, the apparent activation energies for self-diffusion of calcium, silicon, and aluminum are not significantly different from 70 kcal per mole of diffusate,' in agreement with the postulate that these elements diffuse by the same process. For oxygen self-diffusion E is about 85 kcal per mole,' again in agreement with the idea that oxygen is transported,

Jan 1, 1963

Ahrenholz, Herman William, Jr., Student, Lehigh Univ Bethlehem, Pa. '35 Allen, Carl A., Student, Colorado School of Mines Golden, Colo. '35 Allen, Paul W., Student, Mass. Inst. of Tech Cambridge, Mass. '35 Allen, R. M., Jr., Student, Virginia Polytechnic Inst., "G" Co Blacasburg, Va. '35 Amberger, K. H., Student, Michigan College of Min. & Tech Houghton, Mich. '35 Andersen, Burton Lee, Student, School of Mines, Univ. of Washington Seattle, Wash. '35 Anderson, Joseph S., Student, Univ. of Texas El Paso, Tex. '35 Anderson, R. C., Student, Univ. of Washington Seattle, Wash. '35 Ansel, Gerhard 627 Clyde St., Pittsburgh, Pa. '34 Anutta, Frederick P., Student, Michigan College of Min. & Tech Houghton, Mich. '35 Apsouri, Constantin N., Graduate Student, Univ. of Arizona Tucson, Ariz. '35 Arakelian, Haig Mark, Student, Mass. Inst. of Tech Cambridge, Mass. '34 Archer, George K 1420 Court Pl., Denver, Colo. '34 Argall, George O., Jr., K. E. House Golden, Colo. '35 Argiielles, Alfredo, Student, Univ. of Texas El Paso, Tex. '35 Atwater, Ralph, Student, South Dakota School of Mines Rapid City, S. D. '35 Austin, Robert B., Miner, Page Mine, Federal Min. & Smelt. CoSpokane, Wash. '34

Jan 1, 1934

By G. S. Ansell

RCENTLY, in investigating the high-temperature creep properties of an aluminum, SAP-type alloy, MD 2100, fabricated by Professor F. V. Lenel of Rensselaer Polytechnic Institute, Ansell and Weertman1 found unusually low creep rates. These creep rates are 4 or 5 orders of magnitude lower than is predicted by the slowest creep mechanism based upon dispersion-hardened alloy theory. They ascribed these results to a lack of both normal active dislocation sources and a continuous three-dimensional dislocation network. If one couples these results with the following factors of production and microstructure: 1) The compacted and sintered aluminum powder compact is hot-extruded, with an extrusion ratio of 25:1.2 2) The alyminum oxide particles are approximately 140A units thick, and have a spacing varying from 0.05 to 1.5 1. 3 3) The alloy has plate-shaped grains with minor axes of about 5 X 10 -3 mm and major axes several orders of magnitude greater than this.'

Jan 1, 1960

By Bernard Schneider

The diversity of methods available to engineers and geologists for studying rock masses is a reflection of the fact that, despite the rapid and encouraging progress made over the last few years, there is still no unified and comprehensive method which can be used to predict the over- all behavior of a foundation. Nevertheless, it is agreed that the fundamental characteristic affecting the engineering properties of rock is fissuration, and this paper de- scribes a few new or improved methods of studying fissuration and its influence on major engineering factors such as permeability, deformability, or internal stresses in the rock volume. TESTS ON SAMPLES Conventional tests on samples using thin slices can be improved by staining the slices before cutting with certain ionic dyes which have an affinity for certain types of surface. This shows up the microdiscontinuities. For example (quoted terms are French trade names) : "Turquoise Motylhne JSA" will show up porous alteration of basic feldspars. "Vert brillant SA" is used for acid feldspars. "Fushine ASA" will show up voids in all rocks. A mixture of "Turquoise" and "Rhodamine BSA" will show up zones of alteration in all types of feldspars. With this technique, it is easy to make an optical classification of the different structural types, e.g.: (1) rocks completely free from voids; (2) rocks with microfissures, which in turn can be subdivided into (a) rocks with microfissures secant to the mineral structure, (b) rocks with microfissures parallel to the mineral structure, and (c) rocks with ran-

Jan 1, 1970

By C. S. Finney, John Mitchell

The decline of the beehive coking industry was inevitable, but it had filled the needs and economy of its day. A beehive plant required neither large capital investment to construct nor an elaborate and expensive organization to run. The ovens were built near mines from which large quantities of easily-won coking coal of excellent quality could be taken, and handling and preparation costs were thus at a minimum. The beehive process undoubtedly produced fine metallurgical coke, and low yields were considered to be the price that had to be paid for a superior product. Few could have foreseen that the time would come when lack of satisfactory coking coal would force most of the beehive plants in the Connellsville district, for example, to stay idle; and if there were those like Belden who cried out against the enormous waste which was leading to exhaustion of the country's best coking coals, there were many more to whom conservation was almost the negation of what has since become popularly known as the spirit of free enterprise. As for the recovery of such by-products as tar, light oil, and ammonia compounds, throughout much of the beehive era there was little economic incentive to move away from a tried and trusted carbonization method simply to produce materials for which no great market existed anyway. With the twentieth century came changes that were to bring an end to the predominance of beehive coking. Large new steel-producing corporations were formed whose operations were integrated to include not only the making and marketing of iron or steel but also the mining of coal and ore from their own properties, the quarrying of their own limestone and dolomite, and the production of coke at or near their blast furnaces. As the steel industry expanded so did the geographic center of production move westward. By 1893 it had moved from east-central to western Pennsylvania, and by 1923 was located to the north and center of Ohio. This western movement led, of course, to the utilization of the poorer quality coking coals of Illinois, Indiana and Ohio. These coals could not be carbonized to produce an acceptable metallurgical coke in the beehive oven, but could be so treated in the by-product oven. By World War I the technological and economic limitations of the beehive oven as a coke producer were being widely recognized. After the war the number of beehive ovens in existence dropped steadily to a low of 10,816 in 1938, in which year the industry produced only some 800,000 tons of coke out of a total US production of 32.5 million tons. The demands of the second World War led to the rehabilitation of many ovens which had not been used for years, and in 1941, for the first time since 1929, beehive ovens produced more than 10 pet of the country's total coke output. Production fell off again after 1945, but the war in Korea made it necessary once more to utilize all available carbonizing capacity so that by 1951 there were 20,458 ovens with an annual coke capacity of 13.9 million tons in existence. Since that time the iron and steel industry has expanded and modernized its by-product coking facilities, and by the end of 1958 only 64 pet of the 8682 beehive ovens still left were capable of being operated. Because beehive ovens are cheap and easy to build and can be closed down and started up with no great damage to brickwork or refractory, it is likely that they will always have a place, albeit a minor one, in the coking industry. The future role of the beehive oven would seem to be precisely that predicted forty years ago by R. S. McBride of the US Geological Survey. Writing with considerable prescience, McBride declared: "A by-product coke-oven plant requires an elaborate organization and a large investment per unit of coke produced per day. Operators of such plants cannot afford to close them down and start them up with every minor change in market conditions. It is not altogether a question whether beehive coke or by-product coke can be produced at a lower price at any particular time. Often by-product coke will be produced and sold at less than cost simply in order to maintain an organization and give some measure of financial return upon the large investment, which would otherwise

Jan 1, 1961

By M. E. True, B. L. Stone

To cope with the problems encountered when drilling at greater depths and to reduce the amount of physical effort required on the part of drilling crews in making round trips, a new type of semiautomatic power-operated drilling machinery has been developed which permits round trips to be made without the drill pipe being touched by hand. This equipment consists of hy-draulically operated tongs which perform the stabbing, spinning, and tong-ing operations, and two power-operated racking units mounted in the derrick for carrying the pipe to and from the center of the hole and positioning it on the mat and in the rack. With this equipment all of the operations of the various units are controlled remotely by manipulation of hydraulic valve levers and electric switches. INTRODUCTION As horsepower rises to meet the demand for deeper drilling, it is necessary to increase the weight of surface drilling equipment. Also, the duration of repetitive operations in making round trips to change bits is greatly increased with depth. Harder formations encountered at greater depths reduce bit life rapidly, resulting in more frequent round trips and in some instances more time is consumed in changing bits than in drilling. Since the introduction of the first rotary employed to drill for oil in 18951, there has been relatively little change in the general procedure for making up, breaking out, stabbing, and racking drill pipe in making round trips to change bits. The conventional method of making round trips with the drill pipe requires that drilling crews lend strenuous physical assistance to manually operated tools. Although many new devices have been developed to improve efficiency, increase speed of making round trips, improve safety, and reduce the amount of physical effort required in handling drill pipe, it is considered a vital necessity to substitute automatic or semiautomatic power-operated machinery for many of our more or less manual operations if extremely deep drilling is to be carried out successfully on a production basis in the same manner as our shallower drilling of today. Field tests have been made of remotely controlled power-operated tongs, spinner, stabber, and racking equipment. The tongs, which include a built-in pipe spinner, are mounted on a column, are capable of gripping and supporting the weight of a 90-foot stand of drill pipe, and are provided with means for moving the pipe vertically as well as to and from the center of the hole. The racking equipment consists of two units, one at the level of the conventional monkey board and the other at the first girth on the ladder side of the derrick. Most of the physical effort of making round trips with drill pipe has been eliminated in that all of the operations of tonging, spinning, stabbing, and racking of pipe on the floor are controlled by an operator in a seated position by manipulation of hydraulic and electric valve levers and buttons. The driller is relieved of operating the cathead, leaving only the hoisting and slips to be controlled. Complete round trips are made without the drill pipe being touched by the drilling crews. Although the speed of pulling or running a single stand with the power-operated equipment is approximately the same as that with conventional equipment, it is possible to maintain this rate for long periods, whereas fatigue slows crews working conventionally during round trips, especially under adverse weather. After completing round trips with this equipment, drilling crews are not exhausted, consequently, are able to perform effectively their duties while drilling, which is essential for maximum over-all efficiency. Safety is substantially improved as the derrickman performs his duties from a platform enclosed by rails and the tong lines and spinning chain are eliminated from the derrick floor. The hazards encountered in racking drill pipe on the floor and in stabbing are also eliminated. Figure 1 shows a drawing of a derrick equipped with the power-operated drill'pipe handling equipment.

Jan 1, 1949

By M. E. True, B. L. Stone

To cope with the problems encountered when drilling at greater depths and to reduce the amount of physical effort required on the part of drilling crews in making round trips, a new type of semiautomatic power-operated drilling machinery has been developed which permits round trips to be made without the drill pipe being touched by hand. This equipment consists of hy-draulically operated tongs which perform the stabbing, spinning, and tong-ing operations, and two power-operated racking units mounted in the derrick for carrying the pipe to and from the center of the hole and positioning it on the mat and in the rack. With this equipment all of the operations of the various units are controlled remotely by manipulation of hydraulic valve levers and electric switches. INTRODUCTION As horsepower rises to meet the demand for deeper drilling, it is necessary to increase the weight of surface drilling equipment. Also, the duration of repetitive operations in making round trips to change bits is greatly increased with depth. Harder formations encountered at greater depths reduce bit life rapidly, resulting in more frequent round trips and in some instances more time is consumed in changing bits than in drilling. Since the introduction of the first rotary employed to drill for oil in 18951, there has been relatively little change in the general procedure for making up, breaking out, stabbing, and racking drill pipe in making round trips to change bits. The conventional method of making round trips with the drill pipe requires that drilling crews lend strenuous physical assistance to manually operated tools. Although many new devices have been developed to improve efficiency, increase speed of making round trips, improve safety, and reduce the amount of physical effort required in handling drill pipe, it is considered a vital necessity to substitute automatic or semiautomatic power-operated machinery for many of our more or less manual operations if extremely deep drilling is to be carried out successfully on a production basis in the same manner as our shallower drilling of today. Field tests have been made of remotely controlled power-operated tongs, spinner, stabber, and racking equipment. The tongs, which include a built-in pipe spinner, are mounted on a column, are capable of gripping and supporting the weight of a 90-foot stand of drill pipe, and are provided with means for moving the pipe vertically as well as to and from the center of the hole. The racking equipment consists of two units, one at the level of the conventional monkey board and the other at the first girth on the ladder side of the derrick. Most of the physical effort of making round trips with drill pipe has been eliminated in that all of the operations of tonging, spinning, stabbing, and racking of pipe on the floor are controlled by an operator in a seated position by manipulation of hydraulic and electric valve levers and buttons. The driller is relieved of operating the cathead, leaving only the hoisting and slips to be controlled. Complete round trips are made without the drill pipe being touched by the drilling crews. Although the speed of pulling or running a single stand with the power-operated equipment is approximately the same as that with conventional equipment, it is possible to maintain this rate for long periods, whereas fatigue slows crews working conventionally during round trips, especially under adverse weather. After completing round trips with this equipment, drilling crews are not exhausted, consequently, are able to perform effectively their duties while drilling, which is essential for maximum over-all efficiency. Safety is substantially improved as the derrickman performs his duties from a platform enclosed by rails and the tong lines and spinning chain are eliminated from the derrick floor. The hazards encountered in racking drill pipe on the floor and in stabbing are also eliminated. Figure 1 shows a drawing of a derrick equipped with the power-operated drill'pipe handling equipment.

Jan 1, 1949

By Eugene E. Ayres

The centrifuge has come into general use in the oil fields and in the establishments where lubricating and fuel oils are used. Probably centrifugal machinery will be found more useful in the refinery when the operators are more familiar with'the principles of modern centrifuga-tion and when centrifugal engineers better understand the refinery emulsion problems. It is not generally realized that centrifugal force and gravity have two distinct functions: First, subsidence wherein the suspended globules are brought into contact; second, coalescence (when this is possible) by rupturing the films around the globules. In the case of a water-in-oil emulsion, an exercise of the former function alone would yield: (1) oil, free, or as free as desired, from water; and (2) a concentrated water-in-oil emulsion—concentrated in the sense that it contains a relatively large percentage of the internal phase. An exercise of both functions will yield: (1) water-free oil and (2) oil-free water. Centrifugal force is much better than gravity for subsidence because, in the case of industrial emulsions where the suspended globules are visible under a microscope, subsidence is directly proportional to the force applied. Centrifugal force is only slightly better than gravity for coalescence. A minute trace of a proper reagent will immediately do more to nullify the interfacial surface tension opposing coalescence than an indefinite application of the highest attainable centrifugal force. No mathematical analysis has been made to explain the failure of high centrifugal force to induce a decidedly more rapid and more complete coalescence than is obtained by gravity. The adjustment of the centrifuge rotor depends more on the extent of coalescence that is desired or is possible than on any other factor. One reason for this dependence is to be found in the curious relation between the specific gravity and the hydrostatic pressure of emulsions. The continuous centrifuge works on the U-tube or Florentine flask prin-

Jan 1, 1924

By Carl E. Swartz

ALTHOUGH a number of articles appeared during the war advocating the use of cadmium in lead-tin solders, very little information of value can be found in the literature regarding the properties of solders containing cadmium. In most of the published articles the cadmium seems to have been considered merely as a partial substitute for tin in a tin-lead alloy rather than as an essential constituent of a ternary alloy. The primary object of the substitution of cadmium for tin was to lessen the amount of tin required, at a time when the latter was scarce and high-priced, and thus to produce a cheaper solder. While this object was accomplished, the new properties resulting from a change from a binary lead-tin alloy to a ternary lead-tin-cadmium alloy were not given proper consideration in many cases. For the past few years, the American Smelting & Refining Co. has been marketing a very considerable amount of cadmium in the form of lead-tin-cadmium alloys for use as solders for certain special purposes, and the successful use of these solders suggested that if more extensive knowledge were available concerning the properties of lead-tin-cadmium alloys of various compositions their field of usefulness might be materially extended. The investigation described in the present paper was undertaken to obtain this information and, by means of tests under conditions approaching those of practical use, to determine the suitability of cadmium solders for various purposes. CHARACTERISTICS AND REQUIREMENTS OF SOLDERS IN GENERAL General Physical Properties.-When cast into bars, solder should present a smooth and bright appearance free from oxide inclusions and bubbles. Ductility should be as great as possible without sacrificing tensile strength, and vice versa. Hardness and specific gravity are not of any great importance. Melting Characteristics.-The melting or solidification range is important. This statement applies both to the extent of the range and to the actual temperature at which complete solidification occurs. A rather narrow solidification range is desirable in soft solders in order that solidification may take place quickly. The solder should have high fluidity at temperatures only slightly above its melting range. For a

Jan 1, 1928

By George M. Humphrey

It is indeed a pleasure for me to have this opportunity to discuss with you a subject which I find infinitely interesting and intriguing - namely, world-wide iron ore supply and demand. We will be examining the following: (i) current iron ore production and requirements; (ii) the major ore producing countries; (iii) the seaborne iron ore trade; (iv) ore grades or types; and (v) future requirements. For those who may not be familiar with the iron ore mining industry, it may be useful to know at the outset that our industry has been plagued with overcapacity for the past seven years. The optimism which prompted substantial expansion of productive capacity in the early to mid-1970's has unfortunately proved to have been unfounded. I shall begin by describing where we are today as well as a little about how we got there. In the latter regard, it's not so much that historical knowledge will solve any of the industry's problems or provide clairvoyance with regard to the future. Rather it's a feeling on my part that some knowledge of the past will provide a useful perspective capable of enhancing our analysis of future trends. At the time of writing, the most recent year for which we have definitive world-wide iron ore production and consumption statistics is 1980. In that year the world consumed approximately 859 million metric tons of iron ore to produce approximately 510 million metric tons of primary iron consisting of 502 million tons of pig iron and 8 million tons of sponge iron. Crude steel production for the same period attained a level of 719 million metric tons. Because it is possible in today's world to produce steel without consuming iron ore simply by melting scrap, the key statistic insofar as iron ore demand is concerned is primary iron production. In 1980 the ratio of iron ore consumption to primary iron production was 1.68:1.00. The comparable ratio over the past seven years has varied from 1.63 to 1.87 and averaged 1.72:1.00. All the same, as a practical matter crude steel production and iron ore consumption tend to parallel one another closely, and for the time being a good rule of thumb is that to produce an average ton of crude steel requires 1.20 tons of iron ore. Let's turn our attention now to the regions where the 859 million tons of iron ore necessary to support 1980 world-wide iron and steel production were produced. Not surprisingly, most of the ore was produced in the same region where it was smelted. In round numbers 65% of the world's iron ore production is currently consumed in the same general region as it is produced, while 35% enters the seaborne trade and is transported in ocean vessels to a more remote consuming area. Approximately 85% of the world's seaborne iron ore trade is directed to the Western European and Japanese markets, both of which are almost entirely dependent on off-shore sources for their mammoth iron ore requirements. The following tabulation lists, in order, the twelve leading iron ore producing countries in 1980 as well as their percentages of total and free-world production; in addition their export potentials are presented. As a group these countries accounted in that year for 89% of world-wide production which stood at 871 million tons. Leading World Iron Ore Producing Countries [ ] One will first note that world iron ore production is dominated by the top four producers - namely, the U.S.S.R., Brazil, Australia and the U.S.A. - which together accounted for nearly 60% of total 1980 production. Of these four major producers the U.S.A. and the U.S.S.R. produce almost exclusively for consumption in their own regions, whereas Australia and Brazil produce primarily for the seaborne export trade. In fact Australia and Brazil accounted for about 50% of all seaborne iron ore trade in 1980! It is also interesting to note that Australia, Brazil, the U.S.A. and Canada accounted for nearly 60% of free world iron ore production in 1980. Once again the importance of Australia and Brazil shows itself, with their share alone of free world production reaching 35%. As has already been noted, a significant 35% of the world's iron ore production in the year 1980 was

Jan 1, 1982

By Granville S. Borden, Frank H. Madison

The fruits of industry are divided between capital, labor, and governments. Capital takes its redemption and remuneration through profits or dividends; labor takes its share through wages; governments take their share through taxes. Each must have a just share, and the question of what is a just share and the question of who shall make the allocation are vital current political and economic issues. Certain political philosophies con- tend that government should take all the fruits and that some dictator or politician should have irrevocable discretion and power to distribute the fruits. Believers in democracy and free enterprise adhere to the view that capital is entitled to redemption and fair remuneration; that labor shall have a fair wage; and that governments shall make equitable levies. But since all human beings, both individually and in groups, are inherently selfish, there are many variable views in regard to the question of what ratios shall be applied in a division of the fruits and who shall have the power to fix these ratios. Perhaps all parties should devote more energy to means of increasing the crop, and not so much time to quarreling over a division of the existing crop. Moreover, all three parties must guard against killing the goose that lays the mineral eggs. For example, if the tax burden becomes oppressive and the taxpayer refuses to carry it, then business stagnates, with consequent unemployment, reduction of profits, and curtailment of government revenue. If a producer assumes the increased tax burden, marginal operations must be closed down, and, again the consequences are increased unemployment and diminished government revenue. Quite often an increase in a rate of tax results in a reduction of government revenue. Perhaps the worst detrimental feature of selfishness and greed on the part of any one of the three distributees of the fruits of industry is the retardation of expansion of the enterprise. Growth means new jobs, more government revenue, better trade, and better standards of living. Explorations and discoveries have resulted in tremendous contributions

Jan 1, 1964