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Jan 1, 1891

By J. A. Hutcheson

NO one questions that it is technically possible to achieve the controlled release of atomic energy in a form that can be converted into heat or electricity. However, before this is actually an accomplished fact a great amount of work must .be done. This work falls into two general classifications: research and engineering development. At present a research program is under way in this country, co-ordinated and directed by the Atomic Energy Commission, which, as time goes on, will uncover fundamental knowledge of this new science that is essential before industrial applications of atomic energy are practical. Undoubtedly therefore, as the store of available knowledge increases, a time will come when answers to the many questions relating to practicality of atomic energy in industry can be determined with reasonable certainty. That the application of atomic energy in industry will eventually be proved practical is assumed. What then are the engineering problems that must be solved before the first practical electric power generating station can be built?

Jan 1, 1948

By A. C. Fieldner

DURING 1939, 286 by-product coke ovens were completed and put into operation. These included 140 Witputte ovens for the Carnegie-Illinois Steel Corp., at Gary, Ind.; 61 Koppers-Becker ovens for the Ford Motor Co., at River Rouge, Mich.; 65 Semet-Sotvay ovens for the Great Lakes Steel Corp., Zug Island, River Rouge, Mich.; and 20 Curran-Knowtes ovens for the Midwest-Radiant Corp. plant at Mitlstadt, Ill. The Millstadt plant is unique in carbonizing a washed coat containing 3.50 per cent sulphur. The gas contains approximately 2500 grains sulphur per 100 cu. ft.; this is successfully removed with sodium carbonate solution by the Seaboard liquid purification process. The coke is intended primarily for use as a smokeless domestic fuel in the city of St. Louis.

Jan 1, 1941

By AIME AIME

FREDERICK WORTHEN BRADLEY, the newly elected president of the Institute, may be said to be the prototype of the men who have built up the great mining industry of the West. He was born in Nevada County, Cal., on Feb. 21, 1863, and studied mining at the University of California from 1882 to 1884, as a member of the Class of 1885. He was in contact with the mining industry from his earliest years and in his reminiscences which we published last month, he told of meeting Thomas Mein and H. C. Perkins in 1877. Beginning as assayer and roustabout at the Eagle Bird mine, in Nevada county, Mr. Bradley was a deputy mineral surveyor in 1886 and 1887 and superintendent of the Spanish mine from 1885 to 1890. There he made a record of mining gold ore at 65 cents per ton that was not only astounding in that day when modern large scale mining had not yet been developed but is still a mark to shoot at. As a result he was called in 1890 to be assistant manager of the Bunker Hill and Sullivan and was made manager in 1893. In 1897, he was elected to the presidency of the company, a position he still holds. H. C. Perkins and D. O. Mills, the largest stockholders in the Bunker Hill and Sullivan Co. took over the Tacoma smelter in 1898, feeding it with concentrate from the Coeur d'Alene and Douglas Island mines but also doing a gen-era1 smelting business. Mr. Bradley was put in as president of the Smelting company, and held that position until the plant was sold to the American Smelting Securities Co. in 1905.

Jan 1, 1929

By R. G. Lazzell

THE British are worried about the postwar possibilities of their coal mining industry. Indeed, there are causes for this worry, with the aver- age 1943 cost of production at about $5.40 per long ton, compared to $3.20 in 1938. Average production per man day has dropped from 1.12 in 1938 to 1.03 long tons in 1943. As a comparison, production per man day in the United States is approximately 5 long tons. The average day wage in the British mines is now a little aver $4; in 1938 the average day wage was about $2.25. Great Britain produced 227 million long tons in 1938 from their deep mines, and in 1943 production was 194.5 million, with an additional 4.5 million from strip mines. To meet present wartime needs, output is being supplemented by the production of strip- mined coal, an industry that did not exist prior to the war. Strip mining is now producing at the rate of 15 million long tons per year.

Jan 1, 1944

By C. MeCulloch

A NOVEL practice in the bituminous coal industry is the accelerated method of burning coal to ash used in the laboratory of the Fidelity washery of the United Electric Coal Companies, Du Quoin, Ill. Development of this method has supplanted the tedious float-and-sink process for plant control at this plant. In normal operation the plant operators have available within thirty minutes the accurate ash percentage of the coal loaded in the railroad cars and can therefore make adjustments to maintain a practically uniform quality of washed coal, regardless of the variations in washery feed.

Jan 1, 1937

By Verne D. Johnston

ALTHOUGH the stocks of Lake Superior iron ore on dock or at furnaces at the beginning of the year were about 6,000,000 tons less than at the beginning of 1938, the steel industry was operating at only 52 1/2 per cent of capacity and continued at approximately this rate to about Aug. 1. Consequently the iron mines in the Lake Superior district during the winter and spring period were almost all on a curtailed production basis, governed mostly by the needs of the communities rather than by anticipation of ore demands. Many of these properties gave extra employment by prosecuting heavy strip- ping programs, and the underground mines by shaft sinking and rock development. All underground mines re- turned to full operation in September, mostly two-shift, five days per week. A few operated the sixth day. The ship- ping schedules were increased so that the actual delivery for the year exceeded early expectations by about 20 per cent. Final Lake shipments were slightly in excess of 45,000,000 gross tons.

Jan 1, 1940

By Sheldon P. Wimpfen

EVERYWHERE in mining circles the talk is of this new development of hard faced or insert bits which hints of many changes to come in mining practice and rock drill equipment. In the past fifteen years new equipment has appeared in rapid succession which brought one change after another in rock breaking practice. Detachable bits were an innovation which met with wide success. Blasthole diamond drilling has proven itself. Only recently, fusion piercing was announced as a possible solution to the problem of breaking taconite ores. For several years this interest in new methods has been growing and various experimental usages of tungsten carbides in rotary type rock drilling bits have been partially successful. American experts in rock drilling have been concerned with the subject for ten years but actual commercial practice was begun by European manufacturers and mining companies. At this time nearly every manufacturer of rock drilling bits is working on some type of an insert bit. On the Rand, European manufacturers are testing out various types of insert bits. The problems relating to rock drilling in South Africa vary greatly from ours. The bulk of the drilling there is limited to the 31/2-ft holes that are drilled for breaking the gold bearing reef ore. Rock drilling in most mines is far more diversified. Both drifting and ore breaking call for holes varying in length from 4 ft up to as much as 50 ft and more.

Jan 1, 1947

By R. J. Corfield

MODERNIZATION of the entire electrical system and improvement of Rotation process efficiency is the twofold goal of the improvement program underway at the Arthur and Magna concentrators of the Utah Copper Div., Kennecott Copper Corp. Copper ore from the Bingham pit goes to these two concentrators where an interlocking program calls for complete electrical modernization and increased flotation process efficiency through additional capacity and flow scheme revisions. Flotation plants at the two mills are essentially duplicates, and while this article deals with the changes at the Arthur mill similiar improvements are being made at the Magna mill.

Jan 3, 1953

President Howe: It is, of course, not easy to discuss off-hand the paper which Mr. Becker has presented with so much lucidity. I will only make one remark, which is outside of the line of his argument, and concerns merely a passing allusion in his paper. Mr. Becker speaks of the fact that glass breaks with a flat or conchoidal fracture, and that steel shows a rough or " cupped " surface of frac-

Jan 1, 1895

By E. B. REESER

EFFICIENCY is the foundation on which the prosperity of this Nation must be built. Your organization and the members thereof are constantly thinking of Gays and means whereby greater efficiency may be developed. It is often said that this Nation has become great because of its almost up limited supply of raw materials at low production 'costs. This was undoubtedly true in the earlier years but today we face an entirely different situation. We have learned that nature has been very generous in supplying all parts of the world with ample resources of various kinds to meet the demands. This applies to practically all commodities but for the purposes of this discussion, we will limit our remarks to petroleum.

Jan 1, 1931

By Jose J. Alvarez

INTRODUCTION Three major plant processes control the operation of an uranium in-situ mining plant. The three pro- cesses can be categorized under the headings termed - loading, elution, and precipitation. All other plant processes can be subcategorized under the three. In summary, the major plant processes in an in- situ mining plant are: 1. Loading - The pregnant lixiviant from the production wells flows through anion exchange resin. The loading of the uranium on the resin is controlled by various factors, such as ionic strength, valence, functional groups, etc. 2. Elution - The uranium-loaded resin is eluted or regenerated by a high ionic strength solution. The effluent from this process is de- fined as the pregnant eluate. 3 . Precipitation - The pregnant eluate is chemically treated to attain the ideal conditions for the selective precipitation of uranium. Hydrogen peroxide is an ideal precipitating agent under URI conditions. This paper will focus on two of the major processes: Elution and Precipitation. For the most part this relatively young industry has grasped the theory of liquids in continuous flow for the operation of the plant processes. Likewise, URI initially operated the two processes of interest with the continuous flow concept. Through laboratory experiments, URI has abandoned the flow concept and operates the eluting and precipitating stages with the concept of "batch" processes, whereby only one container, filled to a specified liquid level, is chemically processed. In the proceeding sections, each of the two pro- cesses will be discussed in detail, from their original concept to the present operation. Only minor modifications were made to the existing circuits for the conversion with results revealing a greater level of efficiency. ELUTION The elution of the loaded anion exchange resin can be achieved by different high ionic strength solutions. The selection of constituents is dependent on both chemical and economic considerations (i .e., efficiency and operating/restoration factors). The regenerant presently used is a pH 9.5 solution containing 7.5% NaCl and a mixture of Na2C03/NaHC03. expressed as 2.5% NaHC03.

Jan 1, 1980

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

As a result of a changing market pattern for anthracite, there is now an increasing demand for steam size and a decreasing demand for domestic sizes. To help the producer meet this new demand, the authors present the technical aspects of the problem of crushing domestic sizes to steam fineness and size. The market pattern for anthracite coal has changed considerably in recent years. The demand for domestic sizes has been gradually decreasing and the demand for the steam sizes has been increasing. The companies however aim for a maximum production of domestic sizes, since these sizes command premium prices. To meet the market demands, particularly in certain periods of the year, further size reduction of the coarse clean coal is necessary. The present practice of multistage crushing in toothed rolls leads to complicated flow arrangements and is high in cost. Data presented herein indicate that crushers with high reduction ratios could be used with a resulting simplification of crushing flows and an appreciable reduction of capital and operating costs. The different types of crushing equipment investigated included the rotary-type breaker, the jaw crusher, the impactor-type crusher, and the Aerofall mill. CURRENT PRACTICE The standard sizes of anthracite are presented in Table I and market specifications require that any given size be maintained with only a limited amount of oversize and undersize material. The 1958 prices of the various sizes also are presented, so it may be seen that size reduction is only performed when dictated by market demands. At the present time the general practice of crushing anthracite consists of removing the top size and crushing it in a toothed double roll crusher to the next largest size. Consequently, to reduce the domestic egg, stove, and nut sizes to steam sizes of pea, buck, rice, and barley, the original material must be screened to produce the three top sizes and each size must be crushed in three separate rolls to the next largest size. Each crushed product must then be rescreened to remove the oversize for re crush ing. Thus a typical operation would consist of scalping the egg size from a mixture, crushing it to stove size, and rescreening the product. This same cycle would be repeated for the stove and nut sizes requiring three screens and three crushers as well as conveying equipment to handle the crushed products. This results in an expensive series of operations to make a product which sells for a lower price. If the three top sizes are crushed in one operation to the smaller sizes then the yield of steam sizes is low and the production of Nos. 4 and 5 sizes is large. These sizes sell for lower prices than the steam sizes. A typical operation is presented in Fig. 1. From the figure it may be seen that the original cleaned anthracite contains 67.4 pet domestic sizes, 26.4 pct steam sizes, and 6.2 pct fines. After multistage crushing of the domestic to steam sizes the composition of the product is 0.6 pct domestic sizes, 89.6 pct steam sizes, and 9.8 pct fines. From a technical standpoint these results are reasonable but the handling and crushing costs are

Jan 1, 1961

By J. E. Harris

By elimination of other possible nucleation processes, it has been demonstrated, for commercially pure magnesium and a Mg-Al alloy, that at stresses less than that necessary for triple-point cracking creep cavities are most probably nucleated heterogeneously at second-phase particles. Evidence for the existence of such particles is presented and the role of grain boundary sliding in the nucleation process is examined quantitatively. DURING creep deformation of polycrystalline magnesium and its dilute alloys in the temperature range 0.45 to 0.7 Tm (Tm = absolute melting temperature) cavities form mainly along grain boundaries oriented normal to the stress axis and failure eventually occurs by their growth and impingement.' The possibility of low-ductility creep failure of magnesium-alloy reactor cans during service has prompted a number of studies of cavity formation in these materials.2-5 In this paper divergent theories of cavity nucleation6-10 are examined critically and conclusions drawn regarding the most probable mechanism of formation of cavities in magnesium. The paper is divided into two parts. The first section consists of the review and quantitative assessment of the possible role of grain boundary sliding in the nucleation process. In the second part an experimental program is described where an electron-microscope technique has been used to determine if second-phase particles exist in as-received commercially pure magnesium and in a magnesium alloy 1) THEORETICAL REVIEW Possible nucleation sites for creep cavities are: a) high-angle grain boundary triple points,11-14 b) subboundary/grain boundary intersections,15-18 c) ledges in the grain boundary, 19-25 d) second-phase particles. 9,26-30 Each of these possibilities is now considered with special reference to deformation of magnesium. 1.1) Triple-Point Cracking. The stroh31 condition to nucleate fracture is: " > vr [1] where ss = applied shear stress, ? = surface energy, C = shear modulus, L = length of sliding interface. This relationship is independent of the nature of the interface and it is possible to calculate the stress concentration at the end of a sliding high-angle grain boundary and thus deduce, approximately,* the minimum applied stress necessary to nucleate triple-point cracking. Such theoretical predictions for Nimonic alloys have been made by McLean13 and found to agree quite closely with the minimum stress determined by experiment. Stresses capable of nucleating triple-point cracking should also be sufficient to nucleate fracture at a grain boundary due to intragranular slip, for the lengths of sliding interface in the two cases are approximately equal. Thus the minimum stress for triple-point cracking should represent a demarcation between failure due to general "cracking" and that due to cavitation. Hauser, Landon, and Dorn32 have demonstrated that for polycrystalline magnesium in the temperature range 78° to 298°K the fracture stress is inversely proportional to the square root of the mean grain diameter. As pointed out by stroh31 this result indicates that Eq. [I] describes the fracture characteristics of magnesium at low temperatures. Heal3 has reported that for rapid deformation (112 pct hr-1) of polycrystalline magnesium the transition from cracking to cavitation occurs in the temperature range 175" to 275°C. He also presented the fracture-stress values for the same material tested under similar conditions. From Eq. [I] taking G = 1.3 x 10" dynes , ? = 600 ergs . cm-2,33 and L = 0.017 cm, a Stroh-McLean13,31 tensile stress of -266 kg. cm-2 is calculated, corresponding to a transition temperature of -190°C; this is in

Jan 1, 1965

Jan 1, 1905

By John C. Holtz

Hostility is a characteristic of the environment in underground mines. Nature opposes man's efforts to remove mineral deposits, and this condition is recognized when mining is described as winning minerals from the earth. Evidence of this conflict is the action of natural forces to close excavated passageways. In many mines, this behavior of the surrounding environment is an ever-present and ever-changing problem despite much research and empirical study. Geological conditions change abruptly and may result in rock bursts, the inrush of water, or variations in rock hardness. Although aspects of these problems will become intensified as the rate of mining increases, they are not considered in this discussion. Emphasis herein will be on the changes in the atmospheric environment. Men underground work in an ever-changing, alien, ambient atmosphere. Many conditions are encountered that will affect their health and safety unless properly controlled. Potentially explosive gases are released from minerals as they are mined or enter from the surrounding earth. Some mining methods generate noxious gases. Natural processes and diesel equipment can steal life-sustaining oxygen. Airborne dust may cause lung diseases when breathed by miners in excessive amounts and by affecting visibility, introduce safety hazards. Deep mines pose heat problems that may be aggravated by high-speed mining. The deleterious effects caused by these changes, and their mitigation, are subjects of this discussion. THE UNDERGROUND ATMOSPHERE Health and safety problems from changes in the underground atmosphere have been apparent since the beginning of mining. Alleviation of these problems has been by ventilation with normal air from the earth's surface that flushes away the vitiated mine atmosphere. The objective is to provide atmospheric conditions essentially similar to those that man is accustomed to on the surface.

Jan 1, 1970

By B. F. Tillson, C. M. Haight

I. General Remarks..........................723 1. Location............................723 2. Characteristics of the Orebody..................725 (a) Mineralogical (b) Shape, Strike, Dip, Size, Etc. (1) Ore (2) Country Rock 3. Historical Sketch........................725 11. Mining Method..........................726 1. Outline of Methods Now in Operation............... 726 (a) Stopes (b) Fill . (c) Pillars (d) Slices (e) Transportation (f) Miscellaneous 111. Mining Operations........................727 1. Shaft..............................727 (a) Sinking (b) Dimensions and Equipment, Sketches (1) Size, Inclination. Direction, Etc. (2) Timbering (3) Stations (4) Track (5) Hoisting Equipment (a) Skips (b) Engines and Ropes (c) Idlers (d) Man Cars (6) Loading Pockets (a) Operating Figures (7) Headframe

Jan 1, 1918

By Ray W. Carpenter, Ralph Hultgren, Raymond L. Orr

Heats of formation of Au-Zn alloys of compositions Au3Zn and AuZn were rneasured at several temperatures by liquid tin solution calorimetry. The data for Au3Zn show that much smaller heat and entropy effects are associated with the a2- a1 transformation than for the a1 - a transformation. This result is consistent with reported X-ray diffraction studies which indicate that the a1 and (a2 phases have ordered structures which only slightly differ, whereas the a phase is disordered. The heat of formation of AuZn(ß') does not change significantly between 322o and 8000 K, confirming that the ß' alloy remains highly ordered to high temperatures. Not many measurements have been made of the energies of order-disorder transformations in alloy phases. In many cases these energies may be conveniently determined by measuring heats of formation as a function of temperature. Many alloy phases approximately obey Kopp's law of additivity of heat capacities; as a consequence their heats of formation do not change with temperature. In ordered alloys, however, a decrease in degree of order with increasing temperature will contribute an endothermic component to the heat of formation. Thus, for ordered phases where other anomalous contributions to heat capacity in the alloy or its pure components are absent, a measured change in the heat of formation may be attributed to the energy of disordering. A portion of the Au-Zn phase diagram1" is shown in Fig. 1. The composition AuZn(ß') has a well-ordered superlattice structure of the CsCl type at room temperature. The alloy is considered to remain ordered to its melting point,1 although, from X-ray diffraction data and electromotive force measurements of zinc activities, Terpilowski3 concluded that considerable disordering began at -673°K. The Au3Zn composition is a disordered fee solid solution (a) at high temperatures. Between 693o and 543oK, the ailoy orders to form the a1 phase, having a structure based on the L12(AuCu3) type, but with a regular distribution of ordering faults in the c direction, making it tetragonal. Below 543oK, the alloy orders further by means of slight atomic shifts, without diffusion, yielding a complex long-period orthorhom-bic superlattice structure, a2.4-7 Heats of formation of Au3Zn(a2) and AuZn(ß') were measured at 363°K by Biltz et al.8 by aqueous solution calorimetry. Heat capacities of several alloys between 14.9 and 38.4 at. pet Zn in the range from 373" to 723°K were measured by Iwasaki ef al.6 Stoichiometric Au3Zn, however, was not included in the alloys measured, the nearest composition being 26.1 at, pet Zn. Also, the heating rate used for the measurements, 2 deg per min, may not have permitted equilibrium to be reached in the neighborhood of the transformations. Heat contents of solid and liquid AuZn were measured between 488o and 1138°K by Kubaschewski.9 It was decided to determine the ordering energies of Au3Zn and to check on the possible disordering of AuZn by measuring the heats of formation of these alloys at several temperatures. The present paper reports measurements by liquid tin solution calorimetry of the heats of formation of the three phases of Au3Zn at temperatures within their regions of stability, and of AuZn at a low (322°K) and a high (800°K) temperature. EXPERIMENTAL Alloy Preparation and X-Ray Examination. Ten-gram ingots of Au3Zn and AuZn were prepared by melting weighed amounts of gold (99.95 pct Au) and zinc (99.99 pet Zn) together in sealed evacuated Vycor tubes at 1060°K, followed by rapid quenching. The resulting ingots were homogenized at 900°K for 10 days. The Au3Zn alloy lost no weight in preparation, the loss from the AuZn alloy was enough to introduce an uncertainty of only 0.1 at. pet in its composition. Filings for X-ray examination were taken from various parts of each ingot and strain-annealed at 723°K for 20 min in vacuo. The Au3Zn filings were quenched rapidly to retain the disordered phase. Diffraction patterns were determined with an X-ray dif-fractometer using copper Ka radiation.

Jan 1, 1968

By G. L. Smit

The Buffelsfontein Gold Mining Company is situated in the Klerksdorp district of the Province of the Transvaal in the Republic of South Africa, approximately 100 miles west of Johannesburg. Gold production commenced in January 1957 and Uranium treatment of residue slimes as a by-product started in July 1957. The mining property is 5435 mining claims in extent, the equivalent of 7988 acres or 12.5 sq miles. The mine was initially exploited by the Pioneer Twin Shaft System, consisting of a 26-ft diam down-cast shaft and a 22 ft-diam up-cast shaft, sunk in the north-west area of the mine to cater for the hoisting and ventilation requirements for a milling rate of 150,000 tons per month. To exploit the Eastern zone of the property and assist in increasing production to 200,000 tons milled per month, it was decided to sink a further Twin Shaft System consisting of a 28-ft diam downcast shaft and a ventilation up-cast shaft -22-ft in diam, both monolithic concrete lined — 200 ft apart and 5300 ft deep. This paper deals with the sinking and commissioning of these Eastern Twin Shafts. This was achieved by one sinking crew within a period of 20 months. A) PLANNING 1) Siting: The Eastern Shaft System was sited in the north-east area of the property, 12,000 ft east on strike from the Pioneer System. (Fig. 1). Initially the system was sunk to the 5300 ft elevation to conform with the original Pioneer Shaft System. On completion a sub-vertical main shaft will be sunk to the 8000 ft elevation in the immediate vicinity, and subsequently the surface ventilation shaft will be deepened to this elevation by raising from level to level and sliping to full size. The exact positions of the shafts were chosen after: a) underground development had located certain known water bearing zones, and b) geophysical and topographical surveys had been carried out. The shafts passed through the following geological succession: Surface to 4430 ft - Dolomites of the Transvaal System 4430 to 4470 ft - Quartzites and conglomerates of the Black Reef Series of the Transvaal System. 4470 to 4470 ft - Diabasic Lava of the Venters-dorp System (Ventersdorp Contact Reef at base). 4470 to 5250 ft - Quartzites of the Elsburg Series of the Upper Witwaters-rand System. The shafts were sited to ensure that the least possible water would be encountered during sinking, and this contributed to the rapid rate of sinking. 2) Duties: a) The 22-ft diam ventilation shaft was designed to up-cast 1,400,000 cfm of air (0.060 density) against a resistance of 22 in. water gauge. b) The main shaft was designed to handle 160,000 tons hoisted per month as well as the men, material and down-cast ventilation to serve this tonnage. The 284 diam main shaft is equipped with streamlined buntons spaced at 15 ft centers set in pockets which were left during sinking. There are six compartments, four of which are sewed by two dual purpose hoists, hoisting either, twedeck man cages or 13-ton bottom discharge skips. The other two compartments are served by a single purpose winder hoisting double deck cages accommodating 60 men per deck. (Fig. 2). 3) Timing: Permanent equipment was ordered well in advance to insure continuity in the operation. Development underground reached the shaft positions before the start of sinking operations. From this development, boreholes were drilled into shaft positions to drain off water in the vicinity of the shafts. To ensure complete co-ordination of effort, all relevant data, as well as the allocation of responsibility, were made known to all persons and departments concerned. 4) Models: The extensive use of scale models during the preliminary planning stages proved most useful in deciding on design such as buntons, bunton pockets and guide cleats, and complex layouts concerning loading, spillage and pumping arrangements. (Fig. 3). These models also proved invaluable in determining the walling routine. Normal working conditions could

Jan 1, 1964

By F. N. Rhines, A. S. Nemy

The mode of high-temperature tertiary creep of 523-0 aluminum alloy was found to be strongly stress dependent. The occurrence of necking and/or fissures during tertiary creep exhibited a sequence with varying stress, wherein fissures with no necking occur at a minimum stress and necking with no fissures occur at a maximum stress. At an intermediate stress, tertiary creep begins without necking but with fissures in a region just beneath the external surface, followed by necking and the development of fissures along the axis of the specimen. The results are interpreted in terms of grain boundary gliding and the hydrostatic stresses produced by restraints to grain boundary gliding. THERE appear to be two possible reasons for the existence of a period of accelerating elongation prior to rupture during high-temperature creep in metals; namely, 1) a change in the structure of the metal itself, leading to a change in its response to loading, and 2) a reduction in the cross section of the metal resulting in a higher unit loading. Howe,' by annealing and retesting broken low-temperature creep specimens, found that irreparable damage is done once a metal is subjected to creep. Andrade2 first pointed out the association between the reduction in cross section and the increase in creep rate. The present research was designed to investigate both of these factors by retesting specimens machined from material which had been subjected to high-temperature creep testing short of rupture. EXPERIMENTAL METHOD An aluminum alloy, 52s-0, was selected for these studies because it is constitutionally simple, is strong enough to be relatively insensitive to mechanical damage during machining,3 and undergoes accelerating creep in a convenient temperature range. This alloy is hardened by 2.6 wt pct Mg in solid solution and is not subject to overaging effects; it has a very small quantity of an insoluble impurity phase, Mg2Si, finely dispersed throughout its micro-structure. Creep specimens, similar in design to ASTM 0.252-in. tensile specimens, were machined from 3/4-in. round stock and the surfaces were electro-polished. A 24-hr anneal at 500°C was then applied to stabilize the grain size to an average diameier of 0.27 mm. Subsequent heating for as long as 700 hr at the intended test temperatures produced no change in the tensile properties, indicating a well-stabilized state of the material. Corresponding to each of the typical testing conditions listed in Table I, one creep specimen was tested to destruction and this test was employed as a guide in the selection of times for subsequent tests. Duplicate creep tests were then run to various points along the creep curve to produce specimens which were used to determine changes in mechanical properties and microstructures. The latter were examined at the midsections of the specimens. Other specimens were tested to various stages of tertiary creep, after which they were remachined to new cylindrical test specimens, electropolished and re-tested in creep. RESULTS The major findings of this investigation may be illustrated by a comparison of the results of three sets of creep tests. These were run at 400°C at three stress levels, namely 750, 1000, and 1500 psi. In presenting these results in Figs. 1, 5, and 8, the creep curves are given in pairs: That curve of each pair which is drawn through filled circles corresponds to the initial test, and the curve drawn through open circles (and designated by - R) represents the result of retesting the same specimen after machining to a new cylindrical surface. The load was readjusted to the original unit stress at the beginning of each retest.

Jan 1, 1960