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By S. R. Maloof

In 1945, Hollomon' showed that after plastic yielding and prior to necking under simple tension, both ferrous and nonferrous materials are approximated by an equation of the following form: where, a = true stress K = strength coefficient 5 = true strain n = strain hardening exponent If a material is stress free, a plot of log o vs log 6 would be expected to be a straight line, the slope of which is the strain-hardening exponent (n) and the intercept on the ordinate axis for 6 = 1. is the logarithm of the strength coefficient K. It can be easily shown that the strain-hardening exponent in the above equation is equal to the total uniform elongation prior to necking. Thus, one can determine the total uniform elongation prior to necking from such a plot and, as pointed out by C. W. MacGregor, 2 it is the amount of strain prior to necking which determines the extent to which a material may be deep drawn. In the view of the current interest in the fabrication of beryllium shapes, it was thought that a knowledge of the strain-hardening exponent of cross-rolled beryllium sheet would be of general interest. Load-strain data after yielding and prior to necking was taken from the work of Greenspan,3 converted to true stress and true-plastic strain, and plotted on log-log paper as shown in Fig. 1. The value of Young's modulus E for beryllium was taken as 43 . l06 psi. The slope of this plot gave a value of 0.212 for the strain-hardening exponent (n) and the intercept on the ordinate axis a value of 105,000 for K. The total uniform elongation is thus 21.2 pct, in excellent agreement with the value estimated by Greenspan (about 20 pct). Furthermore,

Jan 1, 1960

By T. H. Alden

T. H. Alden (General Electric Research Laboratory)—This paper as well as earlier ones of Dr. Wood represent an important contribution to the experimental description of fatigue fracture. The mechanism of fracture proposed by the authors, however, is not established by this data nor supported by other data existing in the literature. Although taper section metallography provides a rather detailed picture of fatigue crack geometry, photographs so obtained must be interpreted with care. The narrow bands revealed by etching, frequently associated with surface notches, are labeled by the authors "fissures". Measurement shows, taking into account the 20 to 1 taper magnification, that the depth of these structures is at most 2 to 3 times the width. This distinction is important in the conception of a mechanism of crack formation. It is difficult, for example, to imagine a deep, narrow fissure arising from a "ratchet slip" model. A surface notch, on the other hand, may form easily by this mechanism. The notches observed in the present work are the subsurface evidence of the surface slip bands or striations in which fatigue cracks are known to originate.4-6 It is clear that an understanding of the structure of these slip bands is of key importance in understanding the mechanism of fracture. The evidence presented shows that these regions etch preferentially, possibly because they contain a high density of lattice defects, or as the authors state equivalently, because they are "abnormally distorted." However, it is not possible to conclude that the distortion consists of a high density of vacant lattice sites. The fact of a high total shear strain in itself does not assure a predominance of point defects as opposed to other defects, for example, dislocations. Other evidence in the literature which suggests unusual densities of point defects formed by fatigue7-' refers not to the striations or fissures, but to the material between fissures (the "matrix"). If a choice must be made, the preferential etching would seem to be evidence for a high dislocation density, since dislocations are known to encourage chemical attack in copper;g no such effect is known for the case of point defects. A third alternative is that the slip bands are actually cracked, but that near its tip the crack is too narrow to be detected by the authors' metal-lographic technique. In this case the rapid etching can be readily understood in terms of the increased chemical activity of surface atoms. Unless a vacancy mechanism is operative, the motion of dislocations to-and-fro on single slip planes will not lead to crack growth. Point defect or dislocation loop generation are the principal non-reversible effects predicted by this model. In any case, the nonuniform roughening of the surface in a slip band6 requires a flexibility of dislocation motion which is not a part of the to-and-fro fine slip idea. The same is probably true of crack growth by a shear mechanism. Either some dislocations must change their slip planes near the end of the band and return on different planes,'0 or dislocations of opposite sign annihilate." The mechanism by which these processes occur in copper at room temperature or below is that of cross slip. Thus cross slip appears to be essential to fatigue crack growth.6'10"12 The fact that a tensile stress opens the slip bands into broad cracks does not indicate the structure of the bands or the mechanism by which cracks form. The charactersitic concentration of slip into bands during fatigue shows a low resistance to shear strain in these regions. (This fact in itself may be inconsistent with a high concentration of vacancies.) The authors contend also that continuing shear produces an additional mechanical weakening so that the bands fracture easily (are pulled apart) under the influence of the superimposed tensile stress. It is equally possible that the only weakness is a weakness in shear, that the crack propagates by a shear mechanism, and that subsequently the tensile stress pulls the crack apart. Even the direct observation of bands opened by a tensile stress would not be conclusive since, as argued above, they may be fine cracks. The same argument applies to internal cracks, their existence in the presence of a tensile stress not indicating the mechanism of formation. Internal cracks originating in regions of heavy shear have also been seen following tensile deformation of OFHC copper,13 so that this mode of fracture is not unique to combined tensile and fatigue straining. The authors point out in their companion report14 that 90 pct of the cracks formed during pure tor-sional strain were within 8 deg of the normal to the specimen axis. If the tensile stress were an important factor in crack propagation, it is surprising that the cracks cluster about the plane in which the normal stress vanishes. Similarly, a study of zinc single crystals showed that for various orientations the life correlated well with the resolved shear stress on the basal plane,'= and was not dependent on the normal stress across this plane. W. A. Wood and H. M. Bendler (Authors' reply) -Dr. Alden's discussion emphasizes the essential point in the relation of slip band structure to

Jan 1, 1963

By M. J. Pool, with Appendix by Larry Kaufman, R. W. Krenzer

Thermodynmic properties of the Cu-Mn system have been deterrrzined in the temperature range 973" to 1273°K by measuring the vapor pressure of manganese in equilibrium with alloys of compositions varying from 0.13 to 0.91 rnole fraction manganese using the Knudsen effusion technique. Activities of manganese calculated from the measured vapor pressures show large positive deviations. Calculated heats and entropies of mixing are positive with a maximum occurring at approximately 0. 7 mole fraction manganese. ThE investigation of thermodynamic properties of binary systems involving transition metals is of considerable interest at the present time. The incomplete subvalency electron levels of transition metals lead to unusual behavior in some alloys, and considerable information can be obtained by determining the thermodynamic characteristics of these systems and relating the thermodynamic behavior to observed physical properties. The Cu-Mn system, shown in Fig. 1,' provides an example of an alloy system in which the determination of the equilibrium phase diagram and various physical properties has received considerable attention, but very little work has been done on measurement of the thermodynamic properties. Elemental copper and manganese have comparable electronegatives, atomic sizes, and cohesive energies, but differ in crystal structure. They form a continuous series of solid solutions only at high temperatures and the structure varies in a continuous manner from fcc at the copper-rich end to fct at the manganese-rich end.' The purpose of this investigation was to measure the vapor pressure of manganese over a series of Cu-Mn alloys and from these values determine the thermodynamic properties of the binary system. Vapor pressure measurements were made in the temperature range 973" to 1273"K, remaining within the solid-solution region for each alloy studied. EXPERIMENTAL PROCEDURE The relative thermodynamic properties of solid Cu-Mn alloys were determined from vapor pressure measurements by means of the Knudsen effusion tech- nique. Several discussions of the theory can be found in the literature. In the case of copper and manganese, it is assumed that manganese is the only component subliming, so the vapor pressure of manganese over the various alloys can be determined directly by measuring the weight loss of the effusion cell. This is a valid assumption since the vapor pressure of manganese is about four orders of magnitude higher than the vapor pressure of copper in the temperature region of interest. The vapor pressure of manganese over the alloys at any given temperature is then related to the weight loss of the effusion cell per unit area per unit time. An upper limit of 1.5 x 10"s has been measured for the ratio of polyatomic to monoatomic molecules for manganese vapor in equilibrium with solid manganese at 1180°K, so it was assumed that manganese vapor is monoatomic. The effusion cells, fabricated from high-purity alumina, are 26 mm high, 18 mm OD, and have a wall thickness of 1 mm. Three different sized holes of 2.03, 2.54, and 3.18 mm diam were drilled in the lids from the inside face to minimize chipping, and the lids were then slip-cast onto the cells to provide a gas-tight seal. Alumina was selected for this particular application both because of its resistance to thermal shock and reaction with manganese vapor. In the ideal case, the orifice of the effusion cell should have a knife edge so that effusion through the orifice

Jan 1, 1970

By H. William Ahrenholz

The coal industry had another prosperous year in 1966. Since the turn of the decade, production has been climbing at an average rate of 6% per annum. Although the fast pace slackened somewhat, 1966 production showed a gain of almost 3 % over that of 1965. An impressive record can be drawn from the fact that output per man-day in 1966 was more than twice that of the early 1950's, and more than three times that of the early 1940's. And even though man-power requirements lessened, those men retained by the industry worked longer hours and benefited through higher hourly pay rates. However, cost to the customer increased a little. Such statements can only be made for a healthy industry -one that is constantly improving its efficiency. Coal currently provides more than 25% of the total U.S. energy output. About 50% of all U.S.- mined bituminous coal goes directly to the electric utility market. It should be noted that, now, well over one trillion kwh of electricity is generated annually. Dependence of the giant electric utility industry on coal is based on economy and an abundant, dependable supply of this fuel.

Jan 2, 1967

By Alfred H. Bell

The upswing in oil production and drilling activity in Illinois that began in 1937 gained momentum in 1938 and promises to bring a new and higher peak in the state's annual production in 1939. In 1938 the production totaled 23,929,000 bbl., as compared with 7,426,000 bbl. in 1937, more than a threefold increase. The number of producing oil wells in the new fields was 230 at the end of 1937 and it increased to 2157 at the end of 1938. Daily production for the whole state increased from approximately 35,000 bbl. at the end of 1937 to approximately 135,000 bbl. at the end of 1938, nearly a fourfold increase. Of a total of 2539 wells completed in 1938 in Illinois, 1984 produced oil, 26 produced gas and 529 were dry holes. Of the total, 377 are classified as "wildcat" wells, defined as wells drilled outside of proved territory and more than one mile from the nearest production. The remainder, 2162 wells, were drilled in or near proved fields. Of the 377 wildcat wells (Table 2) 32 were successful in discovering oil or gas in commercial quantities, either new fields or extensions of old fields. Four of these discovery wells mere gas wells, which are not yet commercially productive owing to lack of pipe-line facilities. A special effort was made to ascertain the reasons for the locations of as many as possible of the wildcat wells and the results of this investigation are set forth in the following table:

Jan 1, 1939

By Alfred H. Bell

The upswing in oil production and drilling activity in Illinois that began in 1937 gained momentum in 1938 and promises to bring a new and higher peak in the state's annual production in 1939. In 1938 the production totaled 23,929,000 bbl., as compared with 7,426,000 bbl. in 1937, more than a threefold increase. The number of producing oil wells in the new fields was 230 at the end of 1937 and it increased to 2157 at the end of 1938. Daily production for the whole state increased from approximately 35,000 bbl. at the end of 1937 to approximately 135,000 bbl. at the end of 1938, nearly a fourfold increase. Of a total of 2539 wells completed in 1938 in Illinois, 1984 produced oil, 26 produced gas and 529 were dry holes. Of the total, 377 are classified as "wildcat" wells, defined as wells drilled outside of proved territory and more than one mile from the nearest production. The remainder, 2162 wells, were drilled in or near proved fields. Of the 377 wildcat wells (Table 2) 32 were successful in discovering oil or gas in commercial quantities, either new fields or extensions of old fields. Four of these discovery wells mere gas wells, which are not yet commercially productive owing to lack of pipe-line facilities. A special effort was made to ascertain the reasons for the locations of as many as possible of the wildcat wells and the results of this investigation are set forth in the following table:

Jan 1, 1939

By L. E. Young

LOADING machines may be classified in several ways: (1) Machines which cut or break down and load .the coal; (2) machines which simply load the coal; (3) devices which load and transport the coal; (4) de-vices which elevate; (5) devices which transport the coal after it has been shoveled onto the device. There are many types of mechanical loading devices, Bull. 17, Coal Mining Investigations, under the auspices of Carnegie Institute' of Technology, U. S. Bureau of Mines, and the Advisory Board, listing 31 exclusive. of -the large shovels used in thick coal seams. Mechanical loading of coal is not in its infancy; I think some of the early patents go back to 1890. One large coal company, properly called the early leader in this field of mechani-zation, has about 26 large machines at work. This company spent about thirty years in developing a ma-chine, and at the present time is loading approximately two million tons of coal a year in its mines. The chair-man directed a question at the previous speaker; inquir-ing if real progress was being made in mechanical load-ing. The experience of the company just referred to is in itself an adequate answer. A review of a list of mines using mechanical loading devices shows that there are 28 mines which we may say are 100 per cent mechanical. There are four more mines which have been 100 per cent mechanical, but which are idle at this time on account of labor condi-tions. There are at least 25 more at which the tonnage is more than 50 per cent mechanical. In speaking of a 100 per cent mine I do not mean that every pit car of coal is loaded by machine, because when complete, re-covery of the coal is attempted it is not practical to load all the coal at, the break line. In order to load the last 5 to 10 per cent of coal with a minimum of gob material and in order to keep the. working places along the break line timbered in a safe manner it is generally desirable to use a few hand loaders. At a certain mine operat-ing on the block system 90 percent of the coal is being loaded by what is properly a mechanical loading: device, and about 10 per cent -is being loaded by hand. This 10 percent is right off the break line, out of what might be called, the screen pillar. The mines referred to are scattered over the country, . in seven different' States, and the equipment used is produced by ten different machinery companies. As to size, these mines vary from small mines of 500 tons mining capacity-up. In one instance a company with one loading machine is producing an, average of 450 tons per day. One of the large mines that is idle pro-duced at the. rate of 4300 tons per day, of which 97 per cent was actually loaded by machinery.

Jan 8, 1928

By R. A. Deju, R. B. Bhappu

The basic structural unit of all silicate minerals is a tetrahedron with a silicon atom at the center and four oxygen atoms at the corners. The oxygen-silicon distance is about 1.6 & and the oxygen-oxygen distance about 2.6 &. The different types of oxygen-silicon frameworks in the various silicates are based entirely on the combinations of the tetra-hedral oxygen-silicon groups through sharing of oxygen atoms. In recent years, infrared spectroscopy has permitted estimation of the ratio of the ionic character of the oxygen-silicon bond to be a few times that of the oxygen-carbon bond and the ratio of electro-negativity of the two bonds has also substantiated this finding. On this basis, it is generally accepted that the oxygen-silicon bond is the strongest one occurring in silicate minerals. As the oxygen-silicon ratio increases from quartz to the olivines, a greater percentage of the oxygen bonding power is available for bonding to cations other than silicon. Hence, with an increasing oxygen-silicon ratio, there is increasing oxygen-to-metal bonding. Upon the fracturing of a silicate mineral crystal, the oxygen-metal bond, which is almost entirely ionic in character, should break more easily than the stronger oxygen-silicon bond, resulting in a greater number of unsatisfied negative forces on the surface. If, then the mineral is immersed in a liquid containing hydrogen ions, these negative forces should tend to be neutralized by hydrogen ions from solution, resulting in a change of the pH of the solution. An increase in the degree of adsorption of hydrogen ions is to be expected as the oxygen-silicon ratio in the crystal structure increases. This study attempted to correlate the oxygen-silicon ratio for various representative silicate minerals to the adsorption of hydrogen ions. Experiments to determine the effect of surface area on this adsorption and to investigate the effect of surface iron were also conducted. It is believed that such studies of the surface properties of silicate minerals will supply pertinent information about their behavior in froth flotation systems. EXPERIMENTAL PROCEDURE Ion Exchange: Carefully selected, hand-picked samples of the minerals investigated were dry-crushed in a ceramic ball mill and sieved to obtain a minus 100 plus 200-mesh sample. The sized samples were then run through the electro-magnetic separator to remove any iron-bearing particles. A 10-g sample of each mineral was placed in 100 ml of freshly deionized water previously adjusted to a pH of 3.50 with HCl. The beaker containing the sample and the water was placed under a bell jar from which the air was displaced at once by nitrogen gas from a tank. All experiments were carried out in a nitrogen atmosphere. The sample and solution were stirred gently but continuously during the entire experiment, the stirring speed being kept about the same for all runs. Changes in pH were followed with a Beckman Zeromatic pH meter and recorded on a

Jan 1, 1967

By A. E. Brugger

SOME 2,000 years ago Pliny is supposed to have said, "Out of Africa always something new." It may perhaps even now be news to a great many that the Belgian Congo has in recent years been producing approximately a quarter of a ton of diamonds annually. For that is what the 1 1/4 million carats, the recovery from 40-odd operating mines, and about 20 or 25 per cent of the world's normal consumption, amounts to. In 1915, first exploitation operations produced less than 50,000 carats from two mines. The rapid expansion of operations following the Euro¬pean war meant a most rapid development of a virgin land. The interior is now reached by air in one day from the mouth of the Congo River, where steamers unload mail and passengers from Antwerp. This trip formerly consumed 4 to 6 weeks by a narrow-gage, wood-burning, railroad around the rapids of the lower Congo and Livingston Falls; by river boat to the heads of navigation, Djoko Punda on the Kasai or Luebo on the Lulua; and thence by caravan of native porters over narrow-trails to head¬quarters on the Tshikapa. Hundreds of miles of automobile roads have been constructed over veldt and through forest, bridging streams and ferrying rivers. The railway from Capetown to the Katanga copper district to the East, once two months away by trail is now reached in 2 or 3 days over fair automobile highway. Native "boys" who a few short years ago wore little more than a smile and a string of beads and to whom a wheelbarrow was a complicated piece of machinery, are today driving Fords, tending boilers and engines and operating electrical equipment, or are capable mine workmen, machinists and artisans, and dressed in all manner of cast-off white man's clothing. Even as late as 1920, due to difficulties of securing and transporting machinery during and just after the war, the crudest hand operations were the rule for the excavation, sizing, and concentration of the diamondiferous gravel. All washing and concentration is now mechanized, using steam or electric power, and where operations are sufficiently large to permit competition with native labor costs, excavation and transportation are also accomplished by machinery.

Jan 1, 1932

By J. J. Stokes

Literature on the electrolysis of aluminum from cryolite melts and on the structure of these melts is surveyed critically. Data on density, freezing point, and other properties are reviewed. Theories of the electrolysis are examined in the light of these data. Two theories are presented which account equally well for the observations. ALTHOUGH the production of aluminum is a large-scale industrial process, consuming approximately 3 pet of all of the electric current generated in this country, the mechanism of the Hall Process remains a 70-year-old mystery. It is a remarkable fact that the aluminum industry, which was established largely through research and owes much of its growth to research, still employs as its basic process the one developed by Charles Martin Hall while he was still a young man working in a woodshed. There are, of course, modifications in detail and a tremendous expansion of the size of the units used, but basically the process employs carbon anodes and a carbon-lined iron vessel as a cathode for the electrolysis of alumina dissolved in molten salt. Of the almost infinite variety of salts available, the relatively minor Greenland mineral, cryolite, 3NaF.A1F3, is still the only one which gives good performance in the electrolysis. In the early days, the process was thought of as a simple one. The dissolved alumina dissociated and, during the electrolysis, the aluminum was carried to the cathode where it was discharged and collected in a pool of metal. The oxygen was carried to the carbon anode where it reacted to form primarily carbon dioxide together with some carbon monoxide. About the only virtue that can be claimed for this hypothesis is that of simplicity. The actual mechanisms must be tremendously more complex. This paper is devoted to a rather narrow part of the whole subject, namely, what ions are present in the bath, and by what mechanism the current is carried. The complexity of the literature on this system is demonstrated in Table I, which shows the ions postulated as being present during the process. This system is not easy to investigate. Cryolite melts at about 1000°C, so that all the investigations are at relatively high temperatures. Molten cryolite fumes in air, particularly if excess aluminum fluoride is present. The cryolite changes composition on heating at 1000°C, principally by reaction with moisture in the air, but apparently also by reaction with oxygen. Cryolite reacts with most of the customary materials of construction, and those that are not attacked by cryolite seem inevitably to react with molten aluminum. Graphite is the only common material that is inert to the electrolytic bath. Pure cryolite can be held in platinum. There is hope that some of the newer refractories will stand up in bath, and their development for this purpose would be a major advance in the field. The techniques to be discussed, then, are not the most advanced ones. Rather, they are examples of old standardized methods applied under difficult conditions. The principal methods that shed real light on the mechanism are density, electrical conductivity, freezing-point lowering, transport number, viscosity measurements, and reactivity of bath with carbon dioxide. The first measurement to be discussed is density. This procedure is relatively straightforward. A hollow platinum bob is suspended by fine platinum wire from an analytical balance into the melt being investigated. Density1-4 of the system sodium fluoride-aluminum fluoride is shown in Fig. 1. There is a maximum in the density near the cryolite composition, but actually the maximum is on the NaF-rich side of cryolite, close to the composition NaF - Na3AlF8. Density measurements are also useful in giving a clue to the nature of cryolite-alumina melts.1, 4 On the first addition of alumina, density decreases, Fig. 2, although the density of alumina, either solid or liquid, is higher than that of cryolite. These density results indicate that the systems are not simple. A great deal of effort has been spent in measuring the electrical conductivity of molten cryolite and of cryolite with various additives. Conductivity measurements are quite easy for fused salts that can be measured in a dip cell, but no successful dip cell has been made available yet for cryolite. Recourse is had

Jan 1, 1959

By Sixten Wollmar

A RESUME. of some important points regarding the life of drill steel, which have so far been learned from investigations carried out in the SKF Hofors' mines, Sweden, by L. Nordenfelt, is given below. By examining all fatigue breakages which occurred in the test it has bees conclusively proved that when a rock drill of the self-rotating type is used all break-ages of this kind start from the surface of the hole in the steel. It has further been proved that the rough hole surfaces (i.e., the marks left by the indentations of the sand core), are the starting points for the fatigue cracks and that the corrosive action of the water has .a very detrimental influence on the life of the steel. It is now a well known fact that this corrosive action concentrates at the bottom of the indentations left by the sand. The corrosive action of water has been well described by U. R. Evans in his paper, "Some Recent Work on the Corrosion of Metals," delivered on March 31, 1926, before the Liverpool Engineering Society, which is quoted in part below: "It is well known that, under ordinary conditions, most metals do not corrode in water unless dissolved oxygen is present. The strange feature of corrosion in submerged metals, however, is that the attack does not proceed mainly over those portions of the surface which could be reached readily by oxygen; deterioration takes place at certain points which are relatively, inaccessible to oxygen. When. a metal becomes pitted by corrosion, further attack usually deepens the pits, and in many cases it could be proved that there was no free oxygen at the bottom of the pits. None of the older theories of corrosion explain the fact that, although oxygen is necessary for rapid corrosion to take place, the attack actually occurs at the very points which could not be reached directly by oxygen.

Jan 6, 1928

"Located in Madison River Canyon, about 14 miles by road from Norris, Mont. Built in 1907 by Madison River Power Co.DAM: Rock filled wooden crib structure, 183 ft. long, 34 ft. high or 44 ft. to top of flashboards.SPILLWAY: 140 ft. long, equipped with flashboards 10 ft. high.PIPE: One 10-foot and one 12-foot wood-stave pipe, 7,515 feet long.PRESSURE: One pressure chamber, 32 ft. high, 30 ft. wide and 81 ft.CHAMBER: long, built of concrete.PENSTOCKS: 4 9-foot steel penstocks about 200 ft. long, one for each wheel.WATER WHEELS: 4 double horizontal, center discharge, Sampson special 44-inch turbines, 3,600 h-p., at 100 ft. head. Speed 300 r. p. m„ direct connected to generators.GENERATORS: 4 G. E., W. C., 2,250 k-w., 4,000 volts, 300 r. p. m., 3 phase, 60 cycles, alternators, with exciters mounted on same shaft.TRANSFORMERS: 12 G. E., W. C., 900 k-w., 60 cycles. High tension voltage 23,100, 34,650, 46,200; low tension voltage 4,000, 3,800, 3,600. Now operating at 46,200, delta connected. They can be connected Y for 40,000, 60,000 or 80,000.OPERATION: This is the principal plant of the whole system and carries most of the load."

Jan 1, 1913

By John Birkinbine

Secretary's Note.—In printing this paper in the present volume, the figures given in the pamphlet edition have been brought more nearly up to date, the product of 1899 being in many instances inserted. Iron-Ore Resources. In a paper presented to the Institute in 1897,* I discussed the iron-resources of the United States as then developed, calling attention to some principal sources upon which other countries depended. The following statement will give a general idea of the amounts, as compiled from the latest official records, of ores won, and metal produced, in the nations which are important contributors to the world's output. Unfortunately these are not all contemporaneous, since in some countries the statistics do not appear until after one or more years. The statement covers not only pig-iron and iron-ore, but also manganese-ore —the manganese mined being principally used in the production of spiegeleisen or ferro-manganese, and these alloys being generally included in the pig-iron output. Table I, which epitomizes the present situation, shows respectively the countries and the quantities of iron-ore and man-

Jan 1, 1901

By T. B. Massalski, Horace Pops

Martensitic transformations were observed by optical metallography and electrical-resistivity measurements during cooling at cryogenic temperatures of the bcc pr-phase Au-Zn alloys. The com-position dependence of the MS, MF, AS , and AF temperatures have been determined. It is concluded that the martensitic platelets which grow mainly in the lengthwise direction and which are associated with a very small temperature hysteresis are "thermoelastic" martensite. In some cases the observed structure seems to have no definite crys-tallographic habit or consists of a mixture of- plates and broad transformed areas. THE majority of the bcc /3-phase alloys in systems of copper, silver, and gold with zinc and cadmium are unstable at low temperatures and undergo a transformation upon further cooling into another structure. A large number of such transformations are martensitic. The composition dependence of the Ms temperatures has been determined and reported for CU-~n,' Ag-Cd,' and AU-C~~ alloys. The 50 at. pct ordered Au-Zn p' alloy has been reported by Jan and pearson4 to transform martensitically at 58°K in connection with their study of the de Haas-van Alphen effect The p' phase in the Au-Zn system has the ordered CsC1-type structure. The phase field is shown in Fig. 1: it has the typical V shape which is characteristic of all copper-, silver-, and gold-based binary /3-phase alloys. There appears to be no a priori reason why the transformation should be limited to the equiatomic composition and the availability of a cryogenic stage for optical metallography5 has permitted studies down to the liquid-helium temperature of a series of alloys in this system. EXPERIMENTAL PROCEDURE Predetermined weights of 99.999 pct pure gold and zinc were melted and cast under a partial atmosphere of helium in sealed quartz capsules. Thorough mixing of the molten metals was achieved by vigorous shaking, followed by a rapid quench into a brine solution. All ingots were homogenized at 640°C for 4 days. The loss in weight was usually less than 0.04 wt pct and was assumed to be due to

Jan 1, 1965

By E. E. White

I disagree with Mr. Brinsmade on serveral points, and I will discuss this remards under the same six topics which he has used.

Jan 1, 1915

H. W. Turner, San Francisco, Cal. (communication to the Secretary*): This excellent paper apparently represents the results of extensive observation and experience among the mines of the Mother-Lode, but I strongly question its relegation of the dioritic intrusions to the period of fissuring and ore-deposition. The dioritic rocks referred to are evidently those areas represented by Ransome in the Mother-Lode folio1 as meta-diorites, and, as he says, are merely basic facies of the great granodiorite batholite of the Sierra Nevada. If followed downward I venture to predict that these meta-diorite areas would be found to connect with the great batholite, and to be in fact mere apophyses of it. I see no reason for supposing that these apophyses were intruded into the Mother-Lode slates at a period later than the intrusion of the granodiorite in general. The period of fissuring was subsequent to this intrusion; the basic granodiorite or quartz-diorite areas of the foot-hill region are themselves fissured extensively; and the fissures were filled with quartz, apparently during the same general period of the formation of the Mother-Lode veins. Prichard and Ransome2 both refer to the deposition of quartz in the same lode at different periods resulting from a reopening of the fissures. Some time since, I saw in the stopes of the

Jan 1, 1904

By Victor A. Phillips

A series of strips of a Ni-12.7 at. pct A1 alloy were Prepared containing cubical y'(NisAl) precipitates with edge lengths from 60 to 500A. A particle-free solution-tveated strip was included for cornparison. They weve cold-rolled 95 pct and the effects of particle size on the isochronal (1/2 hr) annealing behavior between 300° and 950°C studied (by hardness and light and electron microscopy). It ulas inferred that the particles deformed with the lnatvix becoming lamellae which remained coherent. Comparison with published data fov pure nickel showed that aluminum greatly re-tavded softening and recrystallization, but it made little difference whether or not particles were present. The presence of pakticles led to a heterogeneous distribution of precipitates after annealing at 700" to 750°C. Recovery was not detected. Recrystallization occurred by the growth of new grains into unrecrys-tallized material. In a previous study by the author,' the growth of Ll2-type ordered yl(Ni3Al) precipitates was followed in Ni-12.7 at. pct A1 alloy as a function of aging at 600" and 700°C. The particles were showo to be cubical in shape in all sizes from 50 to 3000A and remained coherent. This work was used as a guide in preparing the starting structures for the present study of the effect of these particles on the annealing behavior of heavily cold-rolled strip. Another question of present interest was whether dislocation and particle hardening were additive, since the structures before rolling ranged from solution-treated to peak-aged to overaged. Also, precipitation might occur on annealing after cold-rolling. Reference may be made to other papers2"5 for previous work in this relatively unexplored field and only some recent work will be mentioned her:. phillips2 studied the effect of deformable 0 to 590A-diam cobalt particles on a Cu-3.23 pct Co alloy rolled 95 pct and found that the particles, which rolled out into thin lamellae, impeded softening and recrystallization. Tanner and servi3 likewise studied the annealing of cold-swaged Cu-2 pct Co alloy containing 150A-diam particles and found impeding effects. Haessner et a1.,4 on the other hand, found that incoherent 2-p-diam non-deformable particles of B4C (0.04 vol pct) tended to increase the rate of recrystallization of copper rolled up to 95 pct reduction. They attributed this to the formation of new grains at the particle interfaces. Humphreys and artin' found that nondeformable silica particles in copper rolled to 30 pct reduction accelerated recrystallization if the particle spacing was large and retarded it if the spacing was smaller. Haessner et a1 4 also studied a rolled Ni-Cr-A1 alloy; however, the particles of y'(Ni3Al)-type precipitate were not put in before rolling, but separated during the isothermal annealing at 750°C. No previous work appears to have been carried out on the effect of y' (Ni3A1) particles on the annealing of Ni-A1 alloy. Hornbogen and ICreye7 redetermined the solubility c of aluminum in nickel as a function of temperature T and showed that it was given by c = 32.6 exp(-1940/RT). This relation gives aluminum solubilities of 15.1, 14.2, 12.0, and 10.7 at. pct at 1000°, 900°, 700°, and 600°C, respectively. The phase precipitated from the nickel-rich solid solution is fcc y1 (Ni3A1) which has a Cu3Au -type ordered structure8 and remains ordered up to 1000°C.B EXPERIMENTAL PROCEDURE The alloy used was identical with that used before. Chemical analysis showed 6.27 wt pct (12.71 at. pct) Al, the principle impurities being 0.065 pct Fe, 0.022 pct Co, 0.020 pct Cu, and 0.004 pct C. Bar stock of 1 in. diam was cold-swaged to % in. diam, cold-rolled to 0.300-in.-thick strip, and annealed at 900°C in dry hydrogen. It was cold-rolled to 0.100-in. thickness and solution-treated for 1 hr at 1000°C while sealed in a quartz tube in argon, quenching in iced brine with the aid of a device to snap off the nose of the tube. Lineal analysis gave an average grain size of 0.055 mm. Pieces of strip were aged at 700°C in vacuo for 30 min, 51/4 hr, and 1 week to produce nominal average particle widths of 60, 150, and 500A, respectively, as known from the previous work.' The average diamond pyramid hardness was determined. The heat-treated strips were rolled from 0.100 to 0.005 in., a reduction of 95 pct, and the rolled strips stored at about -5°C. Small pieces were annealed within 1 week for 30 min at temperatures from 300° to 950° ±2°C in a horizontal vacuum furnace. Strips were withdrawn into a cooling zone, giving an estimated initial cooling rate from 950°C of about 50°C per sec. Average diamond pyramid hardnesses were determined on a lightly electropolished spot on the surface of each strip using 300-g load. Each point on the softening curves represents a separate annealed specimen. Sections containing the rolling direction were examined by optical metallography. Selected specimens were electrothinned to the center plane' and examined by transmission at 100 kv in a Siemens Elmiskop I electron microscope. It is well-known that changes in the structure tend to occur when a deformed strip is electrothinned below a thickness of a few hundred angstroms, although this is less serious with a material such as nickel which has a high melting point, and also is apt to be less serious when particles are present. Observations were nevertheless confined to thicker regions of the foils with estimated thicknesses over 1000A. No changes were observed due to beam exposure.

Jan 1, 1968

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 Lucien Eaton

THE accuracy with which detachable bits can be sharpened and gauged by grinding has made possible refinements in design that cannot be attained by forging. This applies particularly to gauging, where accuracy is of the greatest importance. Gauge changes of 1/32 in., and even 1/64 in., are now feasible, whereas 1/16 -in. changes are about the limit in accuracy for forged bits. These small changes in gauge are feasible, however, only with certain fundamental principles of design. Within reasonable limits the depth to which a hole can be drilled is dependent primarily upon the loss in gauge, or diameter, of the bit. In drilling there is always a certain amount of "overdrilling"-the bouncing of the bit in the hole and its rotation wear away the walls-so that the actual diameter of the hole is greater than the diameter of the bit. The amount of this overdrilling depends upon the kind of rock being drilled, and is sometimes sufficient to permit the use of the same diameter of bit for the full depth of the hole. If, for instance, a 1 3/4-in. diameter bit overdrills 1/32 in. in diameter, and loses 1/64 in. in gauge in drilling 2 ft. of hole, the bottom of the hole is still 1/64 in, larger than the original diameter of the bit, and a new 1 ¾ -in. bit can be used to drill the next 2 ft. If, however, the overdrill is less than the loss in gauge, the next bit must be of smaller diameter. The first requisite in reducing loss of gauge in drilling is to have the points and corners of the wings of the bit cut circles of the same diameter, thus producing a reaming surface with continuous bearing on the wall of the hole. The greater the reaming surface, the less the gauge loss that is to be expected. If the reaming surface, however, is too large, it may impede the rotation of the bit through excessive friction. This principle was probably first recognized in the design of the Carr bit (Fig. 1), a bit with a single cutting edge of wide angle, a definite reaming edge, and a large center hole. The same idea was followed in designing the double-taper cross bits, which are now standard almost everywhere.

Jan 1, 1939

By Nelson Severinghaus

Mineral fillers have been defined as inert materials that are included in a composition for some useful purpose. Because this definition and the scope of this review are not specific-indeed, exceptions abound-accurate total statistics are unavailable. One source estimates the total production in the United States in 1979 at 11.3 Mt (12.5 million st) valued at $534 million, averaging $49.19 per ton. Included are a broad group of minerals-asbestos, barite, bentonite, ground limestone and dolomite, clays, diatomite, feldspar, micas, silica, talc. Additional references will be made to synthetic mineral-derived fillers such as aluminum hydrate, precipitated calcium carbonate, and pyrogenic and precipitated silicas and silicates. A section on related uses at the end of this chapter covers absorbents and filters. Mineral fillers can be included in compounds to accomplish a variety of final results. Hardness, brittleness, impact strength, compressive strength, softening point, fire resistance, surface texture, electrical conductivity, and a host of other physical characteristics may be modified by the inclusion of fillers. These effects are the result of properties of the fillers, including chemical activity, hardness, particle size, shape and distribution, surface structure, color, density, and refractive index. To facilitate an overview of the mineral filler industry, this chapter will review the general usage of fillers, the inherent characteristics that intensify their utilization, and similarly their manufactured characteristics. Major using industries will then be examined in detail. Specific information on deposits and processing for industrial minerals will be found in the appropriate commodity chapter. All of these same minerals have additional significant uses based on their chemical composition or other characteristics. References will be found in other use chapters and summarized in commodity chapters. The chart in Table 1, updated from the 4th edition of Industrial Minerals and Rocks, illustrates the major mineral fillers (both natural and manufactured) and their more important end uses. Use of Mineral Fillers Archaeologists have found that more than 5000 years ago, bitumen used for mortar and waterproofing contained mineral contaminants that acted as fillers. Since that time man has added mineral products to compounds to improve certain characteristics. Perhaps the most common objective of filler usage is reduction of cost of the end product. Certainly the inclusion of a filler that costs 2¢ per 0.45 kg (lb) will reduce the cost of a plastic product whose resin costs 50¢ per 0.45 kg (lb). But the decision to use a specific filler is not that simple. It may "soak up" binder, making a compound more viscous and perhaps preventing its flow into a mold as an injection molding compound, but similarly will give good anti-sag characteristics to putty, caulking compound, or undercoating. Fillers may be incorporated to modify: 1) Cost 2) Physical characteristics 3) Flow characteristics 4) Fire resistance 5) Density 6) Heat conductivity 7) Color or brightness and opacity 8) Hardness, brittleness, impact strength 9) Deformity, viscosity, softening point 10) Processing problems 11) Electrical conductivity 12) Surface texture 13) Thermal expansion characteristics 14) Abrasion resistance

Jan 1, 1983