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By Joseph M. Jacisin

Fatigue tests were conducted on 2017-T4 alumium in alloy in an u1lrcthig.h vacuum of 2 x 10-lo Torr and in air. The vatio of vacuum-to-air faligue life for this ~malerial varied Jrom 3.5:1 at a strain of 2.5 X 1O-3 ill. pegr in. to 8:l at a strum of 3.3 x X ill. per 171. Sotlze pvelivtlinary data were also obtained on the faligue life of the alumirnum alloy in a dvy-nitrogen atmosphere. A dark deposit observed on the fracture surfaces of the aiv-tested specimens was identified as oxides of tr~angarzese and magnesium This observalion, and the close correlalion betaveen results obtained in vaCuum and dry-nitrogen environtments, tend to confirm that air stvongly influences the fatigue mechanism by a covrosion process. THE fatigue life of certain metals is known to increase when the atmospheric pressure is reduced. Gough and Sopwith1-2 postulated that this is due to a joint chemical and mechanical process operating in air at the surface of the metal with oxygen and water vapor playing significant roles. Wadsworth and Hutchings3 observed that aluminum is mainly affected by the presence of water vapor. Very little information is available on the fatigue properties of structural aluminum alloys in a vacuum environment. This report describes work performed on 2017-T4 aluminum alloy in an ultrahigh vacuum (2 x 10-loTorr) and in air. Some preliminary data are also presented on the fatigue life of the alloy in a dry-nitrogen environment. 1) EXPERIMENT 1.1) Test Material. The material tested was commerciaT 2017-T4 aluminum alloy which, in addition to aluminum, has an approximate composition of 4 pct Cu, 1 pct Fe, 0.8 pct Si, 0.5 pct Mg, 0.5 pct Mn, plus smaller amounts of other elements.* Some typical naturally age to a substantially stable condition. Specimens were then machined from these rods and ground to a No. 8 microfinish. In order to minimize variation in material characteristics, specimens were made from rods procured from the same lot. 1.2) Apparatus. The apparatus and test specimen used in the fatigue tests are shown in Fig. 1. The device utilized a single-cantilever beam with rotating-bending loading. Bending stress was applied to the fatigue test specimen by placing a spacer beneath the bearing support, thereby providing a constant deflec-

Jan 1, 1968

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

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

Jan 1, 1954

By P. C. Claspy, P. H. Smakula

The electro-optic coefficients of LiNbO, and KTao.,5Nboe3,O3 (KTN) have been experimentally determined at the HeNe laser wavelengths of 0.6328, 1.15, and 3.39 µ. The coefficients were calculated fronz the retardation induced by an applied electric field. The retardations were measured by observing one or move transmission maxima and minima as a function of the applied field, and by measuring the retardation directly with a Babinet-Soleil compensator. The values obtained agree quite well with those previo~lsly reported by others. Extinction ratios as high as 3000:l have been observed. Resistivities and Curie temperatures have also been measured. Light-induced optical inhornogeneities in LiNbO, repovted by others have been observed. CONSIDERABLE interest has recently been shown in the electro-optic properties of LiNb03 (Refs. 1 and 2) and KTN.3, 4 These crystals have a number of advantages over such well-known materials as KDP. These advantages include: 1) the infrared transmission of KTN and LiNb03 extends to 5 p whereas that of KDP cuts off at 1.5 p ; 2) driving voltages required for modulators using these materials are considerably lower than those required for KDP. After first discussing some theoretical considerations of the electro-optic effect, we will report the results of measurements of the dc electro-optic coefficients of these materials at wavelengths of 0.6328, 1.15, and 3.39 p. THEORY The generalized equation for the electro-optic effect can be written: Although lattice polarization is the fundamental variable, rather than the applied electric field, we follow the convention of expressing the linear electro-optic effect in terms of the applied field. For LiNbO the quadratic terms are negligible. Since KTN has a center of symmetry the rijk tensor is 0. After applying the usual symmetry considerations and index contractions,* the rijk and gijkl tensors can be expressed as follows: LiNbOaT If field directions are now chosen, the change in refractive index and the angle between the induced optic axes and the crystallographic axes can be found. The induced rotations of the principal axes of the index ellipsoid for fields in principal crystallographic directions are given in Table I. Since KTN is initially isotropic, the principal axes in the zero-field condition are parallel to the crystallographic axes. The induced retardations are given in Table 11. EXPERIMENTAL We obtained the various electro-optic coefficients by measuring the induced retardation as a function of

Jan 1, 1968

By B. M. Harra, F. G. McCutcheon, O&apos

Until about 1930, the universal practice of horizontal-retort zinc smelters in the United States, as far as the writers are aware, was to operate the retort furnaces on a 24-hr cycle; that is, the retorts were charged in the early morning of one day, and in the early morning of the following day the residues were cleaned out, the retorts recharged, and a new cycle started. During the business depression of the early thirties, zinc production fell off to the extent that zinc smelters had idle retort capacity. Rather than continue to run some of the retort furnaces on a 24-hr cycle and shut down the remainder, a few companies decided to try out a longer cycle. Although the output of zinc per retort per day is lower, the longer cycle permits a larger charge per retort and improved zinc recovery. Advantages of the longer cycle were soon evident, and the practice spread rapidly. Various lengths of cycle were used, typical examples being 28-hr, 32-hr, 36-hr, and 48-hr. The 48-hr cycle fits in best with the standard routine of smelter practice that was built up on the basis of the 24-hr cycle; that is, the retort furnace "maneuver" always comes at the same time of day, and can be performed in the cool hours of the early morning. Under some conditions, particularly in some of the older plants, where the retorts are smaller than in newer plants, a 48-hr cycle is longer than necessary, and some intermediate length of cycle may theoretically be most economical. Use of an intermediate cycle, however, means that the maneuver comes at different times on consecutive days, and that sometimes it comes during the hottest part of the day. The work on the retort furnaces is extremely hot and arduous at best, and it is much worse, of course, if it is done during the heat of a summer day. When labor was plentiful and tractable, many plants were able to operate on cycles of intermediate length, but the workmen did not like the working hours. Under present labor conditions, it is hardly practicable to operate with a cycle of intermediate length, so the choice has narrowed down to the alternatives of 24-hr or 48-hr cycles. In any discussion of economics, too comprehensive generalization is not warranted, because local conditions affecting economics vary greatly. The best economic practice at one plant at any given time may be uneconomical at another plant, or even at the same plant at another time. To determine the length of cycle that is the better practice at a given plant at a given time requires a weighing of the advantages and disadvantages of each cycle, taking into consideration all the local conditions. some of the factors that must be considered in making the comparison are discussed briefly in the following paragraphs.

Jan 1, 1949

By F. J. Tuck

The Hayden mill of the Ray Mines Division, Kennecott Copper Corporation, is now equipped with two 9 by 12-ft. rod mills having a daily capacity of 6000 tons of coarse-crushing plant product from the mine. These two mills crush a ¾-in. product to approximately 19 per cent on 14 mesh, with only one pass through the mills. Each rod mill has taken the place of four sets of 42-in. Garfield rolls with their attendant screens, and is making a product for the fine-grinding department comparable to that formerly made by rolls in closed circuit with screens. The rod mills are lined with wave-type liners of white iron, varying in thickness from 2 to 3 in. The shell and end liners last approximately 600 to 700 days. Manganese-steel lifter bars, 3? in. thick, are used, spaced about 11 in. apart, and they give a life of 350 to 450 days. Fig. 1 shows the general design and thickness of the wave-type liners and lifter bars; 1½-in. liner bolts are used to fasten the lifter bars to the shell. The shell liners are held in place by the tapered lifter bars. A ¾-in. wood backing is used between the shell and the liners. The rod-mill scoop has removable manganese-steel lips, which last 250 to 300 days. The scoop proper has shown some wear in spots, and these spots have been protected by rubber bolted over them. Water and milk of lime are added to the elevator feeding the mill, but not enough water to cause a pounding of the rods against the shell. Roughly, the water added amounts to 25 to 33 per cent of the ore. Additional water is added to the mill discharge to facilitate distribution to the classifier bowls in the fine-grinding department. Rods are charged to the mill once every four days, the mill being down about 20 min. for this purpose. Other delays are negligible, and usually are due to a loose liner bolt or possibly power trouble. However, these occasions are extremely rare. In charging the mill, a crane is used

Jan 1, 1939

By F. J. Tuck

The Hayden mill of the Ray Mines Division, Kennecott Copper Corporation, is now equipped with two 9 by 12-ft. rod mills having a daily capacity of 6000 tons of coarse-crushing plant product from the mine. These two mills crush a ¾-in. product to approximately 19 per cent on 14 mesh, with only one pass through the mills. Each rod mill has taken the place of four sets of 42-in. Garfield rolls with their attendant screens, and is making a product for the fine-grinding department comparable to that formerly made by rolls in closed circuit with screens. The rod mills are lined with wave-type liners of white iron, varying in thickness from 2 to 3 in. The shell and end liners last approximately 600 to 700 days. Manganese-steel lifter bars, 3? in. thick, are used, spaced about 11 in. apart, and they give a life of 350 to 450 days. Fig. 1 shows the general design and thickness of the wave-type liners and lifter bars; 1½-in. liner bolts are used to fasten the lifter bars to the shell. The shell liners are held in place by the tapered lifter bars. A ¾-in. wood backing is used between the shell and the liners. The rod-mill scoop has removable manganese-steel lips, which last 250 to 300 days. The scoop proper has shown some wear in spots, and these spots have been protected by rubber bolted over them. Water and milk of lime are added to the elevator feeding the mill, but not enough water to cause a pounding of the rods against the shell. Roughly, the water added amounts to 25 to 33 per cent of the ore. Additional water is added to the mill discharge to facilitate distribution to the classifier bowls in the fine-grinding department. Rods are charged to the mill once every four days, the mill being down about 20 min. for this purpose. Other delays are negligible, and usually are due to a loose liner bolt or possibly power trouble. However, these occasions are extremely rare. In charging the mill, a crane is used

Jan 1, 1939

By F. G. McCutcheon

UNTIL about 1930, the universal practice of horizontal-retort zinc smelters in the United States, as far as the writers are aware, was to operate the retort furnaces on a 24-hr cycle; that is, the retorts were charged in the early morning of one day, and in the early morning of the following day the residues were cleaned out, the retorts recharged, and a new cycle started. During the business depression of the early thirties, zinc production fell off to the extent that zinc smelters had idle retort capacity. Rather than continue to run some of the retort furnaces on a 24-hr cycle and shut down the remainder, a few companies decided to try out a longer cycle. Although the output of zinc per retort per day is lower, the longer cycle permits a larger charge per retort and improved zinc recovery. Advantages of the longer cycle were soon evident, and the practice spread rapidly. Various lengths of cycle were used, typical examples being 28-hr, 32-hr, 36-hr, and 48-hr. The 48-hr cycle fits in best with the standard routine of smelter practice that was built up on the basis of the 24-hr cycle; that is, the retort furnace "maneuver" always comes at the same time of day, and can be performed in the cool hours of the early morning. Under some conditions, particularly in some of the older plants, where the retorts are smaller than in newer plants, a 48-hr cycle is longer than necessary, and some intermediate length of cycle may theoretically be most economical. Use of an intermediate cycle, however, means that the maneuver comes at different times on consecutive days, and that sometimes it comes during the hottest part of the day. The work on the retort furnaces is extremely hot and arduous at best, and it is much worse, of course, if it is done during the heat of a summer day. When labor was plentiful and tractable, many plants were able to operate on cycles of intermediate length, but the workmen did not like the working hours. Under present labor conditions, it is hardly practicable to operate with a cycle of intermediate length, so the choice has narrowed down to the alternatives of 24-hr or 48-hr cycles. In any discussion of economics, too comprehensive generalization is not warranted, because local conditions affecting economics vary greatly. The best economic practice at one plant at any given time may be uneconomical at another plant, or even at the same plant at another time. To determine the length of cycle that is the better practice at a given plant at a given time requires a weighing of the advantages and disadvantages of each cycle, taking into consideration all the local conditions. Some of the factors that must be considered in making the comparison are discussed briefly in the following paragraphs.

Jan 1, 1947

By R. S. Shoemaker, F. W. McQuiston

REFINERY SECURITY "Highgrading, " a polite word for stealing gold and silver, has been a way of life since the metals in native form have been mined or produced. Unfortunately, in the past, highgraders were not always severely prosecuted or even looked upon unfavorably in their communities. Recent high metal prices with attendent world wide publicity, however, has made highgrading or thievery again a serious problem in mines and plants. When a small button of gold the size of the end of ones little finger has a value equal to about one week's wages, the safe-guarding of gold production by enforcing tight security measures becomes a prime responsibility of management. The higher the gold price, coupled with unrestricted owner- ship, the greater the temptation to “get-your-cut”, and diligence by management must become an integral part of plant security. Primary security begins by knowing the expected gold production and the sampling of a gold ore mill feed for this determination has proved not reliable. However, solution flows can be accurately sampled and measured thereby closely predicting the gold recovered each day as precipitate. The progressive record of total ounces of gold contained in the precipitation press, referred to in cyanidation plants as I1solution calll1 predicts the accumulated production. The "callH is an important adjunct to security and extreme caution must be used so that the pregnant solution meter and sample are confined in a security area. It should be recognized that this key sample may be the target for Ifdoctoring" by trusted insiders (assay office employees) as a preliminary step for stealing precipitates and/or metal. Security measures similar to the following were adopted after one operation experienced a sizeable theft of gold precipitate and metal by trusted employees. The refinery area where precipitates are handled and melted to bullion and all other gold products are kept should be completely enclosed with a tight security fence preferably of solid construction. Access to this enclosure should be through one of two electrically controlled gates, one for personnel and another for occasional truck entrance. Admittance should be restricted to essential personnel who are required to sign a register in and out with the date and time. I£ there is a "back gate" to the restricted fenced area, it should remain locked except in an emergency and then opened only by a supervisor. Television monitoring cameras viewing the gate entrance and other sensitive areas connected to a viewing screen located in the general office afford visual observation at all times.

Jan 1, 1975

By R. H. McLean

Velocity, kinetic energy and shear in crossflow beneath three-cone jet bits may influence cleaning of the bottom of the borehole and the teeth of the bit. Laboratory investigation shows that each of these parameters is a function of the diameter of the borehole and the product of the volume rate of flow and velocity through the nozzles (QVn). Increasing QV, or decreasing the diameter of the borehole increases each parameter. These functions provide means of predicting the magnitude of each parameter and of scaling the cleaning forces. INTRODUCTION In drilling operations using conventional jet-type rock bits, the impinging jels create an important flow mechanism. Called crossflow, this flow mechanism originates in the impact area of the jets, spreads across the bottom of the hole and supplies the principal source of energy to clean the teeth of the bit and most of the bottom.' Besides providing the means of cleaning cuttings from the bit and the hole bottom, crossflow may also have other, less direct, effects on the rate of penetration. The shear stress generated on the bottom by crossflow will inIluence the thickness and permeability of any filter cake of mud solids or crushed material which forms on the bottom.' These factors may affect the rate of penetration. A previous pulblication introduced some fundamental concepts of crossflow.' Crossllow was shown to occur in a thin layer adjacent to the bottom, and to cover the bottom completely. The maximum velocity in the crossflow above any position on the bottom was found to be directly proporlional to the square root of the product of the volume rate of flow and velocity through the nozzles—the jct QV,,—and inversely proportional to the diameter of the borehole. This information indicated that the effectiveness of crossllow in scavenging the bottom can be improved by maximizing the jet QV. The investigation reported herein amplifies the definition of crossflow. Complete velocity profile data above a representative position on the bottom are analyzed. These data better illustrate control of the capacity of crossflow to scavenge the bottom, and also relate shear stress on the bottom to known, controllable parameters. The conclusion reached in the previous publication that maximization of the jet QV. produces the maximunl cleaning beneath current jet bits is unchanged by these new data; rather, it is strcngthened. Data presented here show that the kinetic energy flux above a representative position on the bottom is maximized by maximizing the jet QV. The shear stress on the bottom will also be shown to be maximized in the same manner. Since the functions relating the jet QV. to velocity, shear stress and kinetic energy also involve the diameter of the borehole, means of equating, or scaling, these quantities in different sizes of boreholes will be illustrated. EXPERIMENTAL EQUIPMENT AND TECHNIQUE JET BIT MODEL Data were recorded from the same laboratory model as used in the aforementioned investigation of the flow around a jet bit.' The model consisted of a 43/4-in. three-cone, jet-type rock bit in a smooth, flat-bottomed borehole constructed of lucite. The bit had a shape and nozzle placement closely resembling larger three-cone jet bits commonly used in field operations. Fig. 1 illustrates the impact area on the bottom of a jet from this bit. Details of the orientation of the jets may be found in the previous publication. TECHNIOU'E OF MEASUREMENT Measurements of the crossflow were made by inserting a very small Pitot tube through the bottom of the simulatetl boreholc. Extreme thinness of the layer of crossflow necessitated accurate measurements of the height of the Pitot tube above the bottom to achieve close definition of the velocity profile. A cathetometer, which could be read to the nearest 0.005 cm, was used to make this measurement.

Jan 1, 1966

By H. B. Charmbury, D. R. Mitchell

Recovery of fine coal from solids reporting to wash water has become increasingly important. These solids range from about 28 mesh to 0 and constitute 3 to 7 pct by weight of the feed tonnage to a coal preparation plant. In the past they were usually wasted to streams, but present clean stream legislation makes it necessary to remove them from the wash water and, if economically possible, recover the associated coal as a saleable product. Where these solids are not wasted to streams or to slime ponds they are recovered by thickeners and filters. Since filter cake is often high in ash or sulfur, or both, it is not readily marketed or used. Common practice is to mix it with other plant products, generally slack or cleaned run-of-mine. Uniform blending is difficult, not only because the impurity is usually high, but also because during certain operating periods it is possible to get a concentration of this filter cake in a railroad or barge.

Jan 2, 1959

By L. W. Davies

A problem often encountered is the provision of materials which have impurity contents below a certain specified level. This problem is in some cases solved by making use of the segregation of the impurity with respect to the liquid and solid states of the material: an ingot of the material is subjected to directional solidification and the ends of the ingot. which are richer in solutes which respectively raise and lower the melting point of the solvent, are removed. An important extension of this process is the zone-refining process invented by pfann,' in which molten zones are passed successively along the length of an ingot and displace solutes towards the ends. Pfann's process is particularly appropriate in cases for which the segregation coefficient k does not differ greatly from unity, or when the impurity content must be kept extremely low (-=,6 atom fraction); it is widely used, particularly in the purification of materials for use in semiconductor electronic devices. It is the purpose of this paper to investigate theoretically the efficiency of the above refining processes; that is! we determine the most economic method of attaining a given yield of solvent of specified purity, taking into account the variables of the process. One variable of interest is the effective value of the segregation coefficient. which depends on the rate of movement of the liquid-solid interface and on the degree of stirring in the liquid; in addition to this, in the zone-refining process we are also interested in the choice of zone length for most efficient operation. A mathematical determination is made here of the lengths of successive molten zones which give, for an initially uniform distribution, the most efficient removal of a solute of arbitrary distribution coefficient. In order to be able to give a quantitative treatment of efficiency, we first define figures of merit applicable to a given distribution of solute. FIGURE OF MERIT Let an ingot of length L contain a distribution y(x) of solute: we confine our discussion to mono-tonic distributions, in order to avoid inconsistencies in the subsequent analysis. The first moments of the distribution about the origin and x = L are respectively L F = hy{x)dx [I] G = f(L -x)y(x)dx = YoL2-F, [21 where yoL is the total amount of solute present. We shall see that the quantities F and G are useful figures of merit for a discussion of the efficiency of zone-refining processes. In the first place, it is

Jan 1, 1960

By L. W. Davies

A problem often encountered is the provision of materials which have impurity contents below a certain specified level. This problem is in some cases solved by making use of the segregation of the impurity with respect to the liquid and solid states of the material: an ingot of the material is subjected to directional solidification and the ends of the ingot. which are richer in solutes which respectively raise and lower the melting point of the solvent, are removed. An important extension of this process is the zone-refining process invented by Pfann,' in which molten zones are passed successively along the length of an ingot and displace solutes towards the ends. Pfann's process is particularly appropriate in cases for which the segregation coefficient k does not differ greatly from unity, or when the impurity content must be kept extremely low (=10-6 atom fraction); it is widely used, particularly in the purification of materials for use in semiconductor electronic devices. It is the purpose of this paper to investigate theoretically the efficiency of the above refining processes; that is! we determine the most economic method of attaining a given yield of solvent of specified purity, taking into account the variables of the process. One variable of interest is the effective value of the segregation coefficient. which depends on the rate of movement of the liquid-solid interface and on the degree of stirring in the liquid; in addition to this, in the zone-refining process we are also interested in the choice of zone length for most efficient operation. A mathematical determination is made here of the lengths of successive molten zones which give, for an initially uniform distribution, the most efficient removal of a solute of arbitrary distribution coefficient. In order to be able to give a quantitative treatment of efficiency, we first define figures of merit applicable to a given distribution of solute. FIGURE OF MERIT Let an ingot of length L contain a distribution y(x) of solute: we confine our discussion to mono-tonic distributions, in order to avoid inconsistencies in the subsequent analysis. The first moments of the distribution about the origin and x = L are respectively L F = hy{x)dx [I] G = f(L -x)y(x)dx = YoL2-F, [21 where yoL is the total amount of solute present. We shall see that the quantities F and G are useful figures of merit for a discussion of the efficiency of zone-refining processes. In the first place, it is

Jan 1, 1960

By George Sachs

I deeply appreciate the honor of being selected to deliver the twentieth Annual Lecture of the Institute of Metals Division. The subject of my paper is extremely involved and voluminous, therefore I shall discuss only some more recent experimental material and speculate somewhat on the physical and mechanical fundamentals of these changes. The flow and the rupture of a solid material are rather different phenomena. It is therefore surprising that up to a rather advanced stage in metallurgical science these two fundamental phenomena have been simultaneously treated and forced into the general conception of the "failure" of solid material. Only in this century, and particularly during the last 20 years, have the very different physical meanings of flow and rupture been clearly recognized. Several hundred of the more recent publications deal with these subjects.1 LAW OF SIMILARITY The basic values commonly considered responsible for the mechanical behavior of any material are the stresses. From this standpoint both a large and a small article made from the same homogeneous metal and of geometrically similar shapes should behave identically, i.e., they should deform geometrically in a similar manner, if subjected to the same stress values at corresponding points. This "law of similarity" was introduced about 60 years ago simultaneously by Barba2 and Kick3 and was derived from the classical theory of elasticity. As far as the plastic behavior of a metal is concerned, this rule appears to be valid. The conventional properties of large and small tension or compression test bars, machined from larger sections, are identical; while the force required to deform a section is proportional to its cross-sectional area and the work consumed in deforming geometrically similar articles in a geometrically similar manner is proportional to its volume. The general trend of the stress-strain diagram of a metal is independent of the section size (Fig. I). No exceptions from this rule are known. However, if the deformation of a metal is continued until rupture occurs, and the breaking stress or the work consumed is measured, a test bar of large cross section appears to fail before or at lower stresses than does one having a small cross section. Thus, the specific impact energy has been often found to decrease with increasing size of the test specimen, particularly for materials with a limited ductility.4 Another exception from the rule of similarity is offered by fatigue-strength values, which usually decrease with increasing section size. 5,6 The conclusion must be drawn from these experimental facts that the ductility of a homogeneous metal decreases with increas¬ing section size (Fig. x). This effect has been actually confirmed by Docherty,7 who observed that in static notch bending tests large specimens failed earlier and in a more brittle fashion than small specimens ma-

Jan 1, 1941

By AIME AIME

MOST of us are far .from home, and yet our Japanese hosts- have made us feel very much at home. Here in the Orient we engineers are .learning a new meaning for the word "orientation"- hereafter that engineering term will be a synonym for gracious hospitality, and "to orient" will mean to turn a whole country over to the World Engineering Congress. Mere geographic distance is no longer able to separate nations, for the conquest of distance is one of the triumphs of 'modern engineering. We of the Americas are fortunate in the position of our continents: with their shores lapped by the waves of two oceans, we face' two great civilizations, the' Occidental and the Oriental, and thus we can share the intellectual resources of these two older cultures.

Jan 1, 1929

By J. J. Kanter

The choice of a nonferrous metal or alloy for a given Application is frequently predicated upon a consideration of properties other than merely the capacity to withstand stress. When ability to withstand stress is the dominant engineering requisite of the material to be used, economic considerations seldom lead to the choice of nonferrous metals and alloys. In most instances where a nonferrous metal is used it is because of some special quality of either corrosion resistance, electrical properties, thermal properties, antifriction, or fabrication advantage that commends its application. The mechanical properties, as a rule, present an incidental engineering consideration entirely secondary to the special quality deciding its use. Indeed, great ingenuity has often been exhibited by engineers and metallurgists in overcoming the mechanical shortcomings of a nonferrous base material. in order to take advantage of some outstanding property. . With this thought in mind, we might better appreciate some of the problems in stress design. encountered with nonferrous metals and alloys pertinent to a symposium on their creep characteristics. Our considerations must necessarily be far from complete because of the broad range of materials in the nonferrous category for which creep properties assume a technical interest. Creep deformation of these metals can be subject for study at temperatures ranging from ambient to incandescent, depending upon which one is under consideration. Whereas the steel metallurgist thinks of creep as an essentially high-temperature engineering consideration, the same is not necessarily true in the field at hand. The diversity of the creep problems in the nonferrous field is perhaps accountable to some extent for its virginity as a symposium subject. Each research has necessarily been sponsored by some organization or group interested either in promoting the use of a certain nonferrous base or in obtaining data toward the solution of a mechanical difficulty encountered in applying a material indispensable for its special attributes. Perhaps as many techniques of creep testing have been used as there are base metals implied under "nonferrous." Some of the creep problems that have occasioned researches are associated with steam-pressure vessels such as tubes, drums, castings, and forgings of copper-base and nickel-base alloys, cable sheathings of lead and lead-base alloys, electrical transmission lines of copper and aluminum, engine parts die-cast from alloys based on zinc, aluminum, copper— soldered and brazed joints-—as well as numerous duties at extremely high temperatures met with alloys based on cobalt, nickel, tungsten, molybdenum, platinum, and other precious or semiprecious metals. Postwar technology will have the benefit of much creep research on nonferrous metals, which for reasons of security cannot be discussed on this occasion. Suitability OF Metals As already intimated, the bulk of the constructional metal that ordinarily is

Jan 1, 1945

By William P. Blake

Copper Basin in Yavapai county, Arizona Territory, about twenty miles southwest of Prescott, is well named. It is a depressed area, and a region of cupriferous impregnation.* The geologic conditions are simple. The foundation rock is a coarse-grained granite and gneiss, in which soda feldspar predominates. There are also dykes of porphyritic rock, and a large quartz vein containing pyrites. Superimposed on this crystalline foundation, we find heavy beds of mechanically formed rocks, conglomerates, breccias, and sandstones, in horizontal layers cropping along the bed of a creek, and apparently the remnants of a much more extended formation, now denuded and largely carried away by gradual atmospheric erosion. The heavy beds of conglomerate are in many places much broken and tilted up, even standing on edge in large blocks, as if they had been lifted by some great convulsion; but the cause is much more simple, being merely the removal by gradual disintegration and erosion of the softened and decayed granitic rock below. The materials of these sedimentary beds are chiefly fragments of granite, gneiss, and plutonic rocks loosely mingled. They are the chief repositories of the copper-ore, which forms the cementing substance. This copper-ore is the blue and the green carbonate, azurite and malachite, and the ore is so generally spread through the mass of the beds that the blue and the green croppings can be seen at a great distance, particularly after or during a shower of rain, when the colors are extremely brilliant and beautiful. The copper carbonate is not only a matrix and a cementing material for the fragments of rock, but it invests and covers these fragments so that only malachite and azurite are visible. The beds are from three to ten feet or more in thickness, and, although seemingly solid carbonate of copper, rarely contain over 12 or 15 per cent. of that metal. Some portions of the conglomerates are much more highly charged with copper than others, and in places the beds are almost without copper. If a lump of the coarse sandstone

Jan 1, 1889

By H. B. Charmbury, D. R. Mitchell

Recovery of fine coal from solids reporting to wash water has become increasingly important. These solids range from about 28 mesh to 0 and constitute 3 to 7 pct by weight of the feed tonnage to a coal preparation plant. In the past they were usually wasted to streams, but present clean stream legislation makes it necessary to remove them from the wash water and, if economically possible, recover the associated coal as a saleable product. Where these solids are not wasted to streams or to slime ponds they are recovered by thickeners and filters. Since filter cake is often high in ash or sulfur, or both, it is not readily marketed or used. Common practice is to mix it with other plant products, generally slack or cleaned run-of-mine. Uniform blending is difficult, not only because the impurity is usually high, but also because during certain operating periods it is possible to get a concentration of this filter cake in a railroad or barge. At some plants filter cake has been derived from the froth flotation clean-up of Dorr thickener underflow. The operating circuits required for a froth flotation plant make this a comparatively high-cost treatment, and in the last few years experiments have been carried out with bulk oil methods. These also are quite expensive and lend themselves only to special situations. Table I gives some characteristics of thickener underflows for plants in the Pittsburgh area treating Pittsburgh seam coal. It will be noted that there is a concentration of impurities in the -200 mesh sizes.

Jan 1, 1960

By Stewart Murray, Philip Klapwijk

INTRODUCTION Forecasting commodity markets is like predicting winners in a horse race: to get the answers right, one needs both luck and judgement. Commodity analysts bear some resemblance to racing tipsters as they try to convince interested parties that they have an inside track on which metal is going to win the great minerals race of the 1990s. Commodity analysts share some features with the gentlemen of the turf: • They influence other people in how to spend risk capital, • They don't put their own money at risk, and • If their tips consistently fail to deliver, then they end up looking elsewhere for employment. In order to find winners we need to know about the kind of race we are in and the strengths and weaknesses of all the runners. Miners are not interested in a five furlong sprint: their race is a long distance affair -- after all, the lead-times between discovery and start-up generally approach a decade for mines that will provide the staying power which leaders require in this industry. We need to look at the environment in which our commodities will be competing: in racing parlance, the going. Our equivalent is the economic background: just as some horses are at a relative advantage when the going is soft, so some commodities will do better than others if economic growth is sluggish. If we are going to pick the winners for the late 1990s, and as an industry, make the right investment decisions, nothing is more important than being able to forecast (with at least a semblance of accuracy) how and where the world's economy is going to develop in the next decade. In addressing the question of whether minerals are going to be leaders in the 1990s, this paper considers the idea that it will increasingly be the world outside the industrialised OECD that will provide the incremental growth in demand for mineral products. Economic growth in the Developing World is fundamentally different from that in the OECD, and the pattern of demand for materials in relation to economic output is much more favourable than in the OECD. If we are to forecast the future, we need to be able to understand the past, both in the boom times of the so called Golden Sixties and in the lean years that followed them. This should help us to decide whether the future is going to be like one or other of these extremes or, on the other hand, a return to normality. Having forecast the going, we can look at individual competitors. • Which are best placed to take advantage of the economic, social and industrial trends that we can now dimly foresee? • Which can resist the continued onslaught of the new materials and technologies? • Which of them will best weather the economic storms (assuming that recessions have not yet been abolished)? If we can provide answers to these questions, we will be well on our way to picking winners, i.e., the metals and minerals that will reward their owners with growing demand, steady (and high) prices leading to profits beyond the dreams of avarice.

Jan 1, 1990

By F. J. Tuck

THE Hayden mill of the Ray Mines Division, Kennecott Copper Corporation, is now equipped with two 9 by 12-ft. rod mills having a daily capacity of 6000 tons of coarse-crushing plant product from the mine. These two mills crush a ¾ -in. product to approximately 19 per cent on 14 mesh, with only one pass through the mills. Each rod mill has taken the place of four sets of 42-in. Garfield rolls with their attendant screens, and is making a product for the fine-grinding department comparable to that formerly made by rolls in closed circuit with screens. The rod mills are lined with wave-type liners of white iron, varying in thickness from 2 to 3 in. The shell and end liners last approximately 600 to 700 days. Manganese-steel lifter bars, 3 5/8 in. thick, are used, spaced about 11 in. apart, and they give a life of 350 to 450 days. Fig. 1 shows the general design and thickness of the wave-type liners and lifter bars; 1 ½ -in. liner bolts are used to fasten the lifter bars to the shell. The shell liners are held in place by the tapered lifter bars. A 3/4-in. wood backing is used between the shell and the liners. The rod-mill scoop has removable manganese-steel lips, which last 250 to 300 days. The scoop proper has shown some wear in spots, and these spots have been protected by rubber bolted over them. Water and milk of lime are added to the elevator feeding the mill, but not enough water to cause a pounding of the rods against the shell. Roughly, the water added amounts to 25 to 33 per cent of the ore. Additional water is added to the mill discharge to facilitate distribution to the classifier bowls in the fine-grinding department. Rods are charged to the mill once every four days, the mill being down about 20 min. for this purpose. Other delays are negligible, and usually are due to a loose liner bolt or possibly power trouble. However, these occasions are extremely rare. In charging the mill, a crane is used

Jan 1, 1938

By Warren B. Weeks

The year 1940 saw an increase of 20 per cent in oil production over the previous year—compared with a 16 per cent increase the previous year. In all, 25,790,380 bbl. were produced, an increase of 4,414,150 bbl, over 1939 and the largest production figure for south Arkansas since 1928. The older settled production continued to decline. Magnolia accounted for most of the increase, with aid from the other recent deep fields. During the year, 169 wells were drilled, 79 fewer than in the previous year. Of 131 wells drilled in proven fields, 18 were dry. Three of the 38 wildcats were pool openers. The drilling depth record was increased from 9028 to 9550 feet. Trend of Prospecting For the first time since 1936, the search for oil in the deep Smackover limestone slackened perceptibly; g dry wildcats ended in Smackover, 2 in. the underlying Eagle Mills; 7 penetrated Paleozoic beds, while 9 ended in Hosston, 2 in Glen Rose, 2 in Paluxy and 4 in Gulf. Geology The wells listed in Table 2 either resulted in the discovery of new oil pools or, because of their depth or location, afforded impor-tant stratigraphic or structural informa-tion. Fig. I is a general columnar and stratigraphic section showing the relation-ship of the strata within the Comanche series. The nomenclature shown was ac-cepted by the Shreveport Geological Society. It was subsequently published by Imlay,l who indicated probable correlations with other areas and presented evidence for the possible Jurassic age of the under-lying Cotton Valley and Smackover formations. New Fields Pouke.—On June 10, 1940, the Louark Producing Co. completed its No. I Sturgis, sec. I, T. 17 S., R. 27 W., in central Miller County, producing between 3583 and 3607 ft. in the upper part of the Paluxy forma-tion. The well flowed I00 bbl. of 3r.4' gravity oil per day through Ns-in. tubing choke. This discovery resulted from seismograph work. Six other wells were drilled in the area, none of which found production in the original sand. One of these wells is producing from a sand 50 ft. deeper in the section. The others were abandoned. No further development is in prospect. Production is from lenticular sands of the upper Paluxy; 45,556 bbl. had been pro-duced by the end of 1940. There is structural closure against a fault to the northwest. McKanzie.—The Atlantic Refining Com-pany's No. I Bodcaw Lumber Co., sec. 29, T. 17 S., R. 23 W., eastern central La-fayette County, was completed June 8, 1940, as the eighth and farthest west of the Smackover (Reynolds o6lite) limestone pools. The well flowed 10 bbl. an hour of 58.6' gravity distillate through an 1 1/64 Ma-in. tubing choke, with a gas-oil ratio of 8400:I.

Jan 1, 1941