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By Edmund M. Warner

Any city dweller who has walked alone along a remote mine passageway has to be impressed by the eerie silence-the total absence of noise except for one's own breathing and scuffing of boots on the bottom. An underground mine can have tremendous contrasts, from complete silence to the ear piercing clatter of percussion drills working in the mine roof. Noise in underground mining machines is both useful and detrimental. Without characteristic noises that are recognizable to miners, operators would have a more difficult job operating their machines. Noise serves as a warning of approaching machinery, and changes in noise give indication that breakdowns may occur.

Jan 1, 1979

I do not think too much emphasis can be placed upon creation of favorable climate for development of foreign mineral resources to feed the hungry maw of our industrialized nation. We have become dependent, not only upon the ten raw products indicated in Mr. Bancroft's article, but upon 71 mineral substances listed by DMPA as in short supply. Forty-four of these are metals, 9 are chemical minerals and 18 are nonmetallic minerals. Of this list only 8 are produced in any large quantity in the United States, and all are vital in peace-time industry. Even more, they are absolutely essential for defense. This represents our present needs. But if our consumption and requirements are projected over the next 25 years as is done by the President's Materials Policy Commission-The Paley Report-our future outlook becomes appalling. The divergence of consumption over production, even allowing for substitution and conservation, will widen in the future. This means our dependence upon foreign raw materials, mainly minerals, will grow. What are the chief foreign sources of minerals for our future needs? They are chiefly Latin America and the dominions and colonies of the Western European nations. The British Commonwealth countries contain in abundance those minerals which the U. S. lacks. Of 32 of the most critical minerals the Commonwealth is deficient in 7 and lacks only 1; whereas the U. S. is deficient in 18. Strikingly, our deficient minerals are abundant in the Commonwealth and several of their deficiencies are in excess in the U. S. Resources of the two nations complement each other. If Europe should be overrun by an unfriendly power that could control the destinies and resources of the European dominions and colonies, our industrial life would be placed in severe jeopardy. Therefore, I heartily endorse Mr. Bancroft's statements regarding the Point IV program, provided the program is wisely administered and implemented. It is from the underdeveloped countries, as pointed out by Mr. Bancroft, that most of our future mineral supplies will have to come. Many of the underdeveloped source countries are African colonies and dominions of Western European nations. Our present aid to Europe, therefore, is not, as many infer, a purely one-sided affair; it safeguards our future industrial development and security. Development aid directly to the underdeveloped countries through a well administered Point IV program, can do much, I believe, to safeguard our prosperity. I regard it as a wise investment for future U. S. mineral needs. Dr. Alan M. Bateman Department of Geological Sciences Yale University Mr. Howland Bancroft's article, The Broadening Road to Foreign Investment, appeared on page 666 of the July MINING ENGINEERING.

Jan 1, 1952

By M. Deighton, R. N. Parkins

The metals Cd, ,Wg-, Sn, TI, Zn, and Zr reach a peak hardness after a criticfir1 deformation by rolling- and then soften with fwther rolling-, thereby exhibiting wovk softening. Optical metallography on Cd, Mg, and Zn shows that work softening is accompanied by a change in grain size occurring during deformation. The creep of zinc from -1l° to +60°C at stresses in the range of 7.9 to 17.3 kg per sq mm is given by € = e0 -t- Bt + Kt +at6. The third and second rate constcints are related by the equation a = K~ K6 and their stress and temperature dependence can be represented by the equations K = A, . exp - (u, - Bu)/kT. A model based upon the stress activated glide of sub-boundaries is proposed which qualitatively accounts for the metallog-raphic observations. Expressions, which are in reasonable quantitative agreement with the ex-pe~inzental observations, are derived for the creep of zinc. THE term "work softening" has been used previously by Polakowski1 and by Cottrell and stokes2 to describe phenomena where further strain of a deformed material leads to a decrease in flow stress. In both cases, however, the conditions were changed for the second straining. Here, the term "work softening" is intended to refer to a decrease in flow stress after continued straining in the same direction at the same temperature: work softening is the antithesis of "work hardening". Work softening of zinc was reported by chadwick3 and in discussion of that paper Jenkins4 indicated that cadmium also work softens. More recently work softening has been reported5 in two magnesium alloys, -99.5 pct Mg. Chadwick found that the hardness, 0.2 pct Proof Stress and the UTS of electrolytic zinc all increased with progressive cold reductions up to 30 pct and then progressively decreased with further rolling. Gay and Kelly6 used a back-reflection X-ray technique to study the effect of cold rolling zinc and found that although deformations greater than 2 to 5 pct reduction in thickness produced some recrystallized grains, deformations greater than 40 pct caused com-plete spontaneous recrystallization. At deformations greater than 60 pct the material was found to consist solely of recrystallized grains, -20 pm diam, the size of which decreased with increasing reduction and was much less than the initial grain size of the annealed material (-300 pm diam). Similar results were also reported for cadmium, tin, and lead.6 Gay, Hirsch, and Kelly' suggest that these experiments indicate recrys-tallization takes place when the dislocation density exceeds a certain value. However, no measurements of MATERIALS AND EXPERIMENTAL PROCEDURE The purity of the metals used in this work is indicated by the following figures: Zn (99.99); Cd (99.94); Mg (99.95); T1 (99.99). The zirconium was iodide crys tal bar with a probable purity of about 98 pct. The metals were obtained in the form of 0.25 in. thick strip or 0.425 in. diam rod by various fabrication methods and then annealed to ensure complete re-crystallization. Hardness-deformation curves were obtained at room temperature by rolling 4 in. thick strip under conditions which left a surface adequate for diamond pyramid hardness tests immediately after rolling. The hardness was taken as the niean of five impressions made using a 5 kg load and the time elapsing between rolling and making the last impression never exceeded 5 min. The zinc specimens for creep testing were from 3 by 3 by 4 in. cast slabs which were rolled to 0.10 in. thickness starting at 350°C and finishing cold. The resulting strip was cut into pieces 1 in. wide and annealed in batches. With suitable choices of annealing temperature between 100" arid 400°C five different grain sizes varying from 4.54 x 10' to 3.03 x l05 grains per sq cm (530 to 20 um diam) were obtained. Creep tests were done in compression using a sub-press, based on a design after Ford,8 in which the strip is compressed between dies 0.100 in. wide under conditions of plane strain. Since there is no lateral spread of the material, the area of contact between the dies and strip remains constant throughout the test and the application of a constant load, using the load maintaining device of a hydraulic testing machine, resulted in a constant stress. Covering the dies with strips of P.T.F.E. reduced frictional effects to a minimum. The creep strain was obtained by measuring the travel of the crosshead of the testing machine to a sensitivity of 0.1 pct reduction in thickness. The complete subpress assembly was contained in a steel box and for tests above the ambient this was filled with liquid paraffin and heated electrically. Temperatures below the ambient were obtained with a cooling mixture of acetone and solid carbon dioxide in the box. The liquids were stirred and the temperature of the specimen, which was controlled to ±0.5"C dur-ing a test, was measured by a thermocouple placed near the dies. Compression testing of cylindrical specimens was also carried out in the subpress using hardened flat discs separated from the test material by P.T.F.E. sheets which obviated barrelling of the specimens. Various initial strain rates were supplied by the hydraulic testing machine, and the deformation was measured by a clock dial gage resting on the cross-

Jan 1, 1970

By T. E. Leontis

ONLY a limited amount of information has been published on the effect of thorium as an alloying ingredient in magnesium. McDonald13'2 showed that the addition of thorium in amounts up to 3 pct increases significantly both the ductility and the strength of rolled magnesium sheet. A recent publication by Sauerwald" indicates that thorium contributes strength and creep resistance at elevated temperatures to magnesium in both the cast and wrought states. The observations of Sauerwald are confirmed in this paper, which presents the results of a systematic study of the effects of compositional variation on the tensile properties, creep resistance, conductivity, density, and metallography of Mg-Th alloys containing up to 50 pct Th both in the sand-cast and in the extruded states. The effect of the addition of cerium to Mg-Th alloys is also reported. Experimental Methods The alloys studied in this investigation were pre.pared in small laboratory melts according to the melting practice described by Nelson4 as the crucible method. Electrolytic magnesium was used as the starting material to which the thorium* was introduced in metallic form. It was observed that the behavior of thorium in molten magnesium with respect to reaction with protective fluxes is similar to that of rare-earth metals. In order to obtain the highest possible alloying efficiency, it is necessary to use non-MgC1, fluxes and to apply all the precautions described by Marande6 for melting, alloying, and casting magnesium-rare-earth alloys. .Under these conditions it is possible to obtain an alloying efficiency in the range of 85 to 95 pct. The Mg-Th-Ce alloys were made with cerium of the highest purity available, not with the commercial product "misch-metal" which contains approximately 50 pct Ce. The total rare-earth content of this material was 97.1 pct, of which 0.3 pct was rare-earth metals other than cerium; the major impurities comprising the balance were iron and magnesium. The thorium and cerium contents of each alloy were determined chemically. Test bars 6% in. long with a 2% in. long reduced section of % in. diam were cast in sand molds using a four-baf pattern. In most cases, a 3 in. diam extrusion billet 10 in. long was cast in a steel mold from the same melt used for making the sand-cast test bars. For the higher thorium alloys, it was necessary to remelt the foundry scrap from the test bars in order to make the extrusion billets. The billets were scalped to a diameter of 2 15/16 in. and the ends faced to a length of 9% in. as limited by the size of the extrusion container. The alloys were extruded into % in. diam rod on a 500-ton direct-extrusion press. The details of the extrusion process are as follows: Billet preheat 925°F (2 to 16 hr) Container temperature 900 °F Die temperature 900°F Extrusion speed 10 ft per min Reduction ratio 36:1 Percent reduction 97.3 Alloys containing up to 10.2 pct Th were preheated at 925°F for 2 hr; all alloys with higher thorium contents were preheated for 16 hr. In order to establish the effect of extrusion temperature, billets of

Jan 1, 1953

By Notman, Arthur

THIS is a day of surpluses, some good and some not so good. One can hardly pick up a newspaper, magazine, review or economic treatise without confronting the fact that we have or are threatened with more .of certain goods than we have use for, whether it be corn, wheat, cotton, coal or oil, -lumber, 'leather, newsprint, automobiles, zinc, lead, copper or what not. How did this unprecedented state of affairs arise? Who is to blame? How shall we avoid smothering ourselves in our own production? The voice of the farmer is raised in appeal to the Government to relieve him from the dilemma of his own fertility. In answer the Government employs thousands of experts and spends millions of dollars to aid the farmer. But to what end? To make him still more efficient so that he may produce even greater surpluses? Not satisfied with the result, it spends other millions to prepare vast areas of the arid land of the West for more farmers to cultivate and insure yet larger surpluses. Then what? Naturally another appeal to the Government to fix prices by financing the

Jan 1, 1928

By Raymond D. Sloan

Development and exploration in Oklahoma showed a definite increase in 1941 over 1940 in production, wildcat activity and total well completions. Geological and geophysical work continued and possibly increased slightly in amount. Oklahoma retained third place in the ranking of oil-producing states of the nation, topped only by Texas and California, and followed by Illinois in fourth position. Production averaged 416,800 bbl. daily, a 1.7 per cent increase, as compared with 409,900 bbl. daily for the year 1940. Maintenance of the state's relatively stable producing rate resulted from the discovery of numerous small pools, the recombing of old areas, the establishment of new producing horizons, and the drilling of an increased number of marginal wells. The market for the Mid-Continent crude was strengthened materially to meet the increased demand. Coincident with a deeline in production in Illinois, shipment of Oklahoma-Kansas crude to the eastern points averaged 453,000 bbl. per day, an increase of 49.4 per cent as compared with 1940. The movement south amounted to an average of 42,000 bbl. per day, a loss of 41.7 per cent. Secondary-recovery programs in Oklahoma were given more consideration, operators realizing that in the future many pools would necessarily have to resort to this type of operation. Five permits for water-flooding were issued by the Corporation Commission during the year. These projects were in Rogers, Wagoner and Nowata Counties, in the shallow northeastern area, which has relatively extensive development of this type in the Bartles-ville sand at from 600 to 1000 ft. in depth. Gas repressuring of an individual lease in the 35-year old Glenn pool, Creek County, proved relatively successful, and after a year's operation was extended to cover a much larger area, with several operators participating. More of this type of work is expected in the future; however, it was given an initial setback when the legislative proposal authorizing secondary recovery in the Wilcox zone of Oklahoma's largest pool, the Oklahoma City pool, was pigeonholed by the committee. A continuation of the low discovery rates within the state would possibly lead to a revival of interest in the project. Many pools formerly classified as prorated areas were removed from this classification and transferred into the group of settled production. At the close of the year the issuance of the Federal Conservation Order M-68 and the interrelated Preference Rating Order P-98 caused a sharp decline in drilling operations, which is expected to continue until such operations are stabilized at a lower rate. The ultimate effect of these two orders on discoveries, production and reserves is a matter of speculation and conjecture. During the year, 2110 new wells were drilled, an increase of approximately 18 per cent as compared with the previous year. Of the total, 1235 were oil wells with a combined initial potential of 342,754 bbl. daily, 178 were gas wells with a total

Jan 1, 1942

By Raymond D. Sloan

Development and exploration in Oklahoma showed a definite increase in 1941 over 1940 in production, wildcat activity and total well completions. Geological and geophysical work continued and possibly increased slightly in amount. Oklahoma retained third place in the ranking of oil-producing states of the nation, topped only by Texas and California, and followed by Illinois in fourth position. Production averaged 416,800 bbl. daily, a 1.7 per cent increase, as compared with 409,900 bbl. daily for the year 1940. Maintenance of the state's relatively stable producing rate resulted from the discovery of numerous small pools, the recombing of old areas, the establishment of new producing horizons, and the drilling of an increased number of marginal wells. The market for the Mid-Continent crude was strengthened materially to meet the increased demand. Coincident with a deeline in production in Illinois, shipment of Oklahoma-Kansas crude to the eastern points averaged 453,000 bbl. per day, an increase of 49.4 per cent as compared with 1940. The movement south amounted to an average of 42,000 bbl. per day, a loss of 41.7 per cent. Secondary-recovery programs in Oklahoma were given more consideration, operators realizing that in the future many pools would necessarily have to resort to this type of operation. Five permits for water-flooding were issued by the Corporation Commission during the year. These projects were in Rogers, Wagoner and Nowata Counties, in the shallow northeastern area, which has relatively extensive development of this type in the Bartles-ville sand at from 600 to 1000 ft. in depth. Gas repressuring of an individual lease in the 35-year old Glenn pool, Creek County, proved relatively successful, and after a year's operation was extended to cover a much larger area, with several operators participating. More of this type of work is expected in the future; however, it was given an initial setback when the legislative proposal authorizing secondary recovery in the Wilcox zone of Oklahoma's largest pool, the Oklahoma City pool, was pigeonholed by the committee. A continuation of the low discovery rates within the state would possibly lead to a revival of interest in the project. Many pools formerly classified as prorated areas were removed from this classification and transferred into the group of settled production. At the close of the year the issuance of the Federal Conservation Order M-68 and the interrelated Preference Rating Order P-98 caused a sharp decline in drilling operations, which is expected to continue until such operations are stabilized at a lower rate. The ultimate effect of these two orders on discoveries, production and reserves is a matter of speculation and conjecture. During the year, 2110 new wells were drilled, an increase of approximately 18 per cent as compared with the previous year. Of the total, 1235 were oil wells with a combined initial potential of 342,754 bbl. daily, 178 were gas wells with a total

Jan 1, 1942

By Michael A. Zappia

INTRODUCTION The San Manuel mine of Magmacopper Company is located approximately 72 km northeast of Tucson, Arizona. The mine, concentrator, smelter, electrolytic refinery and continuous rod casting are located on the west side of the San Pedro River valley. Ore Production: 58 1W mtpd Production To Date: Over 340 000 000 mt since 1956. Ore Grade: 0.70% net sulfide cu. As conceived at the present time, the San Manuel ore body is the lower half of what was originally an elliptical (In cross sect ion) cylinder same 2400 m long with major and minor cross sectional axes of 1500 m and 760 m. After emplacement, the cylinder was diagonally sliced along its long dimension by the San Manuel fault, which separated it into halves, moving the upper half same 2500 m down dip and bringing a wedge-shaped blanket of tertiary conglomerate into position over the lower segment of the cylinder. The lower segment is the San Manuel ore body. The upper part, moved down dip, is the Kalamazoo ore body. Originally, these were one ore body. The San Manuel ore body is a structurally weak and highly fractured quartz monzonite, the principal ore minerals being chalcopyrite, with small amounts of molybdenite. The present mining method is full gravity diagonal retreat panel caving by block. This is shown on Figure 2 , (Block Sequence 56, 57, 58, 59 --on the following page.) The ore mined is hoisted to the surface for transportation to the concentrator. BLOCK CAVING MINING SEQUENCE HISTORY The San Manuel mine was planned and developed as a full gravity block caving mine. It continues as a full gravity block caving mine. Since its initial start, the major changes in mining have been in the areas of ground support, rate of draw and undercut sequence. Ground weight resulting from block caving plays a dual role. It can cause extensive damage to ex- traction openings, yet it is necessary to break the ore into easily handled size. The ground- weight problem resolves itself into finding ways to make it manageable. The initial early ground support in the production areas consisted of timber sets and various combinations of steel and timber sets. This sup- port system proved unsatisfactory and a new support system utilizing concrete was developed. Concrete support, 30 cm to 45 cm thick, proved satisfactory. Rates of draw were experimented with. A draw rate of 45 cm to 61 cm (32 to 45 mt) per draw point per day was determined as optimum in con- trolling ground weight. Undercut sequence or mining sequence was recognized as a major contributing factor to ground weight. This is the most difficult factor to change. The lead time from primary development to undercut can be as much as two years. Initial Checkerboard Pattern The initial undercut sequence was in a checker- board pattern (Figure 2). Blocks Sequence 1, 2, 3, 4, 5, 6--next page.

Jan 1, 1981

By Reaves. M. J.

The title of my talk, "What Happened to the Uranium Boom?" is old news. Certainly it is for this group. All of us that make our living in uranium know that the boom of the last half of the 1970's is over. U.S. production has been exceeding consumption by more than two to one. Mines and mills are closing and yellowcake prices have been dropping for over 20 months. The gloomy outlook for the industry in the near term has been well documented by soothsayers of various descriptions, your daily newspapers, and in the Nuexco Monthly Reports. I'd like to attempt to describe the next upturn in the market (speculate, really) based upon the clues we're seeing now. In order to do that, I'd first like to go over briefly, some of the market factors that contributed to the recent price drop and resultant production cutbacks, and then hypothesize on the way these factors are changing and will change. Market prices are greatly affected (maybe even entirely determined) by buyer perceptions. This is particularly true with uranium, because of the long lead times associated with nuclear plant construction and also with conventional mine/mill development. Before the price rise (say, 1975) utility uranium buyers believed that: 1) U.S. producers would have difficulty expanding to meet U.S. demand. 2) Australian and Canadian production was essential to avoid shortages in the early 1980's. 3) Uranlum prices would continue to rise as demand exceeded supply. 4) Enrichment capacity would become inadequate. It was thought necessary, therefore, to build enriched inventory in the early 1980's for use in the late 1980's. Artificially accelerated expansion of the uranium producer industry was necessary to accommodate anticipated enrichment demand. Current perceptions are largely the opposite. These are the beliefs that were held most of this year and late last year as prices dropped. 1) U.S. production is far in excess of domestic need. Contraction of the U.S. production lndustry is necessary. 2) Canadian and Australian supply is optional and not essential. Producers in those countries are expanding mainly by displacing higher cost production and not because they fill a void, 3) Prices may be essentially stable for some time. 4) Enriched uranium is in excess supply. That is 1981. 1982 is shaping up to look like this: 1) Prices will have bottomed out. (That is not Nuexco's opinion necessarily, by the way, but it is my opinion.) 2) There will still be substantial utility inventories, but fewer spot sales. 3) Canadian and perhaps Australian sellers will have made substantial sales in the U.S. and will be aggressively seeking more. 4) U.S. production will have been dramatically curtailed. U.S. utilities that wish to con- tract long term will have difficulty in finding domestic sellers. Concern will develop about the availability of U.S. production capability. Virtually all long term con- tracts signed will be with non-U.S. sellers. 5) An awareness will begin to develop among U.S. buyers that we are approaching a period of dependence upon foreign uranium (which will be true). The history of the uranium market has been one of dramatic changes and overreaction to those changes. The rapid price rise of a few years ago generated excess U.S. production capacity and the rapid price drop of the last two years will almost certainly result in too little capacity. It will soon be difficult for U.S. buyers to buy domestic material except on the spot market. The question is, "will they care?" The lack of demand, of course, is the underlying reason for the current poor health of the uranium industry. In 1972, 1973 and 1974 collectively, there were 105 nuclear reactors ordered in the U.S. That ordering rate was expected to continue and accelerate throughout this century. In 1975, 1976, 1977, 1978, 1979, and 1980 altogether, there were 56 more reactors cancelled than ordered. The net growth of our only customer since 1974 has been a negative 56. TO put this in perspective, if these 56 reactors were operating now it would more than double present U.S. uranium consumption. Underlying lack of demand is something that is simply not going to change in this decade. Time is going to be required. The NRC indicates that the maximum feasible number of new reactors that can be licensed each year is six. That would increase uranium consumption by only 10% per year. New reactors, if ordered tomorrow, would not generate new uranium demand until after 1990. Even so, United States' consumption of uranium will rise from the 1980 level of 18 million pounds per year, to

Jan 1, 1982

By R. G. Bowman

Discussions of processes for the manufacture of white lead generally open with the statement that white lead is the oldest chemical pigment known to man. This fact is of more than historical interest; in the light of the present extensive use of white lead, it indicates that the compound possesses characteristics that make it unique among white pigments, and superior to all in its particular field. A greater variety of processes have been proposed for manufacturing white lend than for any other one chemical compound. Only a few of these processes have come into commercial use, however, and a large part of the present production here and abroad is made by the process used in ancient times. The outstanding disadvantages of the old process are that it is slow and laborious, affords no control of the product, and yields at best an impure material lacking in uniformity. Improved processes, sometimes classed all together as "quick processes," have had as their main object the shortening of the time required. Most of them accomplish this and some have yielded a superior product as well, but only two or three are now in use commercially. For centuries the manufacture and mixing of paints have been treated as an art, and the practice has been much influenced by tradition and precedent. There is a reluctance to abandon old materials for this purpose, and as few new uses for white lead have been developed, the adoption of the quick processes has been slow. The electrolytic production of white lead is not new, a large number of processes and forms of apparatus have been proposed and patented within the past 25 years, all directed toward the production of white lead, wholly or in part, by electrolysis. One or two of these processes were tried commercially but proved unsuccessful. Process The Sperry process, used by the Anaconda Lead Products Co., is the invention of Elmer A. Sperry, and was worked out to a commercial basis at the East Chicago plant. It resembles, in a general way, earlier methods, but embodies the features that make it possible to operate

Jan 1, 1926

By Niels Hansen

The substructure formed by drawing at room temperature in aluminum (99.5 and 99.998 pct purity) and in recrystallized aluminum aluminum-oxide products containing from 0.2 to 4.7 wt pct of aluminum -oxide was examined by transmission electron microscopy, and the flow stress of the drawn materials was measured by tensile testing at room temperature. A sub-grain structure was present after a reduction in area by drawing of 10 to 20 pct, and the subgrain size was observed to decrease with increasing deformation. The tensile data show that the increase in flow stress (0.2 pct offset) by drawing from 10 to 95 pct depends on the reduction in area, not on the composition of the materials. Dispersion strengthening and subgrain bowzdary strengtlzening contribute to the flow stress, and these strengthening processes have been found to be linearly additive. The flow stress (0) can be related to the subgrain size &) by the Petclz relation = uo + k . t, where go is dependent on the composition of the products and k is approximately the same far all materials. THE microstructure of dispersion-strengthened aluminum aluminum-oxide products consists of small oxide plates distributed in an aluminum matrix. The matrix structure depends on the manufacturing history, and in hot-worked as well as cold-worked products the matrix is divided by subgrain (or dislocation) boundaries. For hot-extruded products it has been shown1 that dispersion strengthening and subgrain boundary strengthening are linearly additive, and the flow stress (0.2 pct offset.) at room temperature has been related to the subgrain size (ts) by a Petch equation,2,3 s = so + k . ts-1/2, where s0, increases with increasing oxide content. For cold-worked products containing subgrains no systematic work has been reported, and it was the aim of the present study to examine the microstructure and the relationship between the flow stress and the subgrain size for such products. The behavior of' aluminum aluminum-oxide products depends on the purity of the aluminum matrix, and aluminum of the matrix purity (99.5 pct) was included in the investigation. The literature contains few data about the behavior of this impure aluminum, and aluminum of a higher purity (99.998 pct) was therefore also examined. As regards the relationship between the flow stress and the subgrain size in cold-worked dispersion-strengthened products, no systematic work has been reported. For aged cold-worked structures containing fine precipitates (Fe-Mo carbide) a Petch relation has been found,4 and it has been shown that the k value NlELS HANSEN is Head, Metallurgy Department, Danish Atomic Energy Commission, Research Establishment Riso, Denmark. Manuscriot submitted January 9, 1969. IMD is approximately the same as in iron, whereas the s0 value is higher owing to the presence of the precipi-tates. Investigations of metals such as tungsten,' ferrous metals,4,6 and molybdenum7 cold drawn or swaged at room temperature have shown that the flow stress can be related to the subgrain size by a Petch relation when ts is taken as the subgrain size perpendicular to the direction of deformation. For aluminum no work has been reported on the relationship between the flow stress and the subgrain size after deformation at room temperature, whereas for aluminum tensile strained at different temperatures in the range -183" to 375°C a Petch relation has been found by taking ts equal to the subgrain size.' In the present study two aluminum materials (99.998 and 99.5 pct) and three aluminum aluminum-oxide products (containing 0.2, 1.0, and 4.7 wt pct oxide) were drawn at room temperature to reductions in area from about 10 to about 95 pct. The structures were studied by transmission electron microscopy, and the flow stress (0.2 pct offset) was measured at room temperature. EXPERIMENTAL Materials. The materials are given in Table I together with the chemical analysis. The three aluminum aluminum-oxide products were manufactured from aluminum powder that had been compacted and Table I. Chemical Analysis of Materials Al203 Fe SI Material wt pct wt pct wt pct 99.998 pct* - 0.0004 0.0012 Aluminum 99.5 pctt - 0.36 0.16 Aluminum AlMD13† 0.2 0.16 0.12 Aluminum- -Oxide AlMD105† 1.0 0.26 0.18 Products SAPISML960s† 4.7 0.22 0.19 *Other impurities: O.0004 pct max each of Cu and Zn (supplier's analysis). †Other impurities: 0.03 pct max Cu. 0.02 pct max each of Mn, Mg, Zn, Ti. Table 11. Mean Diameter of Aluminum-Oxide Particles in Extruded and in Cold--Drawn Aluminum Aluminum-Oxide Products Mean Diam. of A1203-Plates* Material State A AlMD105 Extruded 540 Cold Drawn 97 pct 510 SAP ISML 960 Extruded 770 Cold Drawn 95 pct 820 *The standard deviation of the mean is approx. 5 ±pct.

Jan 1, 1970

By H. B. Menardi

THE history, geology, ore deposits and current mining operations of the Huitzuco district have been described by C. W. Vaupell1 and the current mill operations by David Segura.2 This paper completes the story of Huitzuco with a description of the reduction of the flotation concentrates. The predominant ore mineral in the Huitzuco deposit is livingstonite, a rare, mercury-antimony sulphide (HgS.2Sb2S3). Other associated minerals include cinnabar, stibnite and elemental sulphur. The gangue is largely limestone and gypsum. Former operators mined a selected grade of ore, from which the mercury was recovered by an expensive, inefficient operation. A consid-erable tonnage of rejects and fills, carrying mercury and antimony, was left in the mine. In order to recover both mercury and antimony from these fills and also from newly developed ore, a combination method has been developed, by means of which the livingstonite is concentrated by flotation and the flotation concentrates subsequently reduced to metallic mercury and antimony. An essential part of the flotation treatment, as developed by Segura, is the use of copper sulphate as an activating agent. A typical analysis of concentrates now being produced at the property is as follows: Hg, 7.73 per cent; Sb, 26.40; Fe, 1.20; S, 29.68; CaO, 11.50; MgO, 7.25; Si02, 1.63; A1203, 0.65.

Jan 1, 1939

By H. Kenworthy, M. M. Fine

The progress made to date in the mineral dressing of complex Arkansas titanium ores is reported in this paper. Concentrates of rutile, a dioxide of titanium, were produced by treating a submarginal ore from the Magnet Cove Rutile Co., Inc., Magnet, Ark. The rutile recovery at present is 45 to 50 pct of the total TiO2 with the possibility of increased recovery with additional research. The work reported herein is part of a general study which will later include results of investigation of other titanium ore bodies in the Magnet, Ark., area. The fragile character of the rutile minerals, the softness of the gangue, the occurrence of pyrite and leucoxene, and the considerable proportion of extremely small size rutile in the presence of montmorillonite- and illite- type clays in the Magnet Cove open-pit mine are the important factors in the mineral dressing of this submarginal rutile ore.

Jan 1, 1949

By G. E. Warner

This paper presents a review and analysis of a highly stratified Burbank sand waterflooding project in Osage County, Okla. Permeability values in this reservoir range from less than 0.1 md to nearly 3 darcys, and 90 percent of the permeability capacity is contained in approximately 26 percent of the reservoir volume. To date, the project has recovered 103 bbl/flood-acre-ft since waterflood initiation, which represents 50 percent of the total preflood cumulative and 12 percent of the original stock-tank oil in place (OSTOIP). The outstanding feature of this project has been the economical recovery of a sizable quantity of oil at water-oil ratios higher than those normally considered prohibitive. Nearly 70 percent of the recovery to date has been achieved at water-cut values greater than 95 percent. An ultimate waterflood recovery of 138 bbl/flood-acre-ft, or 67 percent of the preflood recovery and 16 percent of the OSTOIP, is predicted. The ultimate primary and secondary recovery will be 302 bbl/acre-ft and approximutely 40 percent of the OSTOIP. The actual performance is compared with the predicted performance by both the Stiles and the Dykstra-Parsons methods, and applicabiliry of these methods to this particular reservoir is evaluated. The performance of this project indicafes highly stratified reservoirs can be successfully and economically waterflooded when either reservoir or economic conditions preclude the use of mechanical and chemical means of mdbility control. Introduction The concept of the stratified reservoir has been used by reservoir engineers for many years to describe the vertical variations in permeability within a reservoir. Numerous techniques have been advanced to predict the expected waterflood performance of the stratified reservoir, and in most cases they prove reasonably accurate. The advisability of undertaking a prospective waterflooding project is determined, based on these predictions. This paper presents a review of a waterflood project in a highly stratified reservoir that normally would be considered a marginal prospect, at best, when evaluated with any prediction technique properly adjusted for reservoir conditions. This kind of project can be successful if the predictions are used to design a waterflooding system capable of economically recovering an oil cut of 1 to 2 percent. Experience gained in the early stages of the project's development was profitably applied in the later development of the entire reservoir and could prove beneficial to other projects with similar reservoir stratification or fracture characteristics. The East Burbank pool waterflooding operations have been developed in three separate stages, each of which has been a separate project: Mid-Burbank Unit, Flood A, and Stanley Waterflood. When referred to collectively, these projects will be called the Composite Project. The Skelly Oil Co.'s Barber lease — the only portion of the pool not included in the three projects — is operated as a separate facility with a cooperative injection-well agreement. It contains less than 80 productive acres and would not significantly affect the over-all performance picture.

Jan 1, 1969

ALEXANDER BRYDEN Alexander Bryden, clean of engineers in the anthracite mining field, and a member of the pioneer Pittston family, was claimed by death Wednesday, September 26, 1917. The announcement was not a surprise, as he had been in a critical condition for several days. Mr. Bryden had been in failing health for a year past, but his indomitable will kept him at his accustomed duties until quite recently. Two weeks before his death he attended the regular weekly meeting of the Board of Directors of the Miners' Savings Bank in Pittston, Pa. At that time he apparently felt that his end was fast approaching, as he bade farewell to his associates on the bank board. Until a few days before the end, he continued to give attention to business matters at his home, in preparation for the end, which he met unflinchingly. In the passing of Mr. Bryden, northeastern Pennsylvania loses one of the best-known mining engineers, and the community, a man who, throughout his long life, was actively interested in its social welfare and business prosperity. He was for many years prominent in the management of the mining operations of the Pennsylvania Coal Co., and at the time of his death held the position of consulting engineer of the Pennsylvania and Hillside Coal. Companies.

Jan 12, 1917

By Robert L. Fleischer

A well known but unexplained experimental fact is the observation1 that aluminum shows a pure <111> wire texture, in contrast with other fcc metals, which have a mixed <100> <111> texture at moderate reductions. A possibly related single crystal observation on aluminum is the following:2 under tensile deformation <111> single crystals are stable (maintain their original axial orientation), while <100> crystals are unstable at room temperature and become stable at lower temperatures. This suggests that at room temperature, as proposed by Yen and Hibbard,3 the single <111> texture obtains be- cause <100> orientations are unstable in the poly-crystal. It also implies that for wire drawing at lower temperature a <100> component of the texture might be expected since that orientation is then stable. This possibility has been examined by wire drawing at temperatures near 77°K. The apparatus was designed to be used with an existing draw-bench, a steel trough acting as die holder and containing sufficient liquid nitrogen to immerse the wire entering the die. The trough is insulated by a styrofoam container to prevent excessive boiling of the bath. The wire emerging from the die enters a polyethylene tube, into the far end of which the grips are inserted. Liquid nitrogen is funneled into the tube to maintain the low tem-

Jan 1, 1962

By Maurice Grieves

The great majority of shafts constructed today are still excavated by drilling and blasting, a method which changed very little in over 100 years until the introduction of the mechanical lashing unit and cactus grab by the South Africans, which enabled muck to be removed as fast as massive hoisting systems could handle it and resulted in very rapid rates of sinking. Record breaking month&apos;s performances were achieved at -Hartebeestfontein No. 4 shaft, October 1960-337.1 m; Western Reefs No. 4 shaft, October 1961-340.7 m; and Buffelsfontein eastern twin shaft, March 1962-381.2m. The method was very labor-intensive, requiring a crew of over 60 workers at the shaft bottom during the drill cycle. Safety precautions were strict, but in the drive to achieve rapid advance, cases of personal injury were still somewhat high because of the large number of people engaged in this potentially hostile environment. The South African method, as it came to be referred to throughout the rest of the world, was adopted in the United Kingdom in the late 1950s in a modified form with greatly reduced manpower and nonsimultaneous sinking and lining, which was insisted on by the British Mines Inspectorate. In that instance, it was successfully used to sink the 7.3-m-diam concrete-lined shafts at Kellingley to 770 m depth, with rates of advance of over 90 m/month achieved, a British record at that time. During the sinking of the 1.15-km-deep twin shafts at Boulby potash mine in the UK in 1970, the method was again used, but for the first time ever in Britain exemptions from the mining code permitted the use of crash beams, crash doors, jack catches, and semi-simultaneous sinking and lining techniques. New British shaft sinking records of over 120 m/month were achieved in both shafts. Similar equipment and techniques were used in the early 1970s to sink several deep shafts in Canada, notably Creighton #7 and the Con zinc mine at Yellowknife in the NW Territories. Today, this equipment is standard for deep shafts in the US and the rest of the world. However, with the tremendous escalation in mining labor costs, the impact of health and safety legislation, and environmental regulations, coupled with a very real shortage of miners willing to work in this exposed situation it was apparent that an alternative to the labor-intensive conventional method of shaft construction had to be found. Recognizing the trend is inevitable, one or two major German shaft sinking contracting firms began to take a fresh look at full face boring techniques applied to tunnels and raise bored shafts. The results were most encouraging. Tunnel drivage techniques using moles had developed considerably from Colonel Beaumont&apos;s original channel tunnel machine circa 1880 to the superbly engineered Priestley machine selected to cut the British side in 1975 and the double shielded Robbins Grandori borer on the French side of the English Channel. Full face tunnel machines were being successfully used to drive uphill in inclined shafts in Austria and Switzerland. At Mapprag in Switzerland, the Demag mole drove the first (intentional) vertical transition and curve, and then went on to successfully complete the 730-m-long penstock shaft at an inclination of 35°. In Austria, the Wirth mole drove the Kaprun Glacier ski-lift railcar tunnel at record breaking rates of 457 m/MONTH (best 30 m/d) through green schist at an inclination of 29° for a distance of 3.35 km while the Hydro tunnel at Sarrelli in Switzerland was being driven by the Robbin&apos;s mole at an inclination of 35°. Simultaneously, extremely promising results were obtained using large assemblies of cutter discs on raise borer heads, such as the 4.87-m-diam X 460-m-deep shaft raised by Teton for Jim Walter Resources Inc. Bearing in mind that most mine shafts in the future (unless in exceptionally competent rock) will require some form of lining, and the trend will be toward deeper shafts as the more easily accessible mineral deposits become exhausted, it was seen that normal raise boring had definite limitations in vertical accuracy, in the limitation imposed by the drill string on the available torque that could be applied to the cutter head, and in the risk of collapse of the unsupported shaft rock wall in friable or jointed and fissured ground, since it is not possible to apply any form of temporary support until the permanent lining is being installed. A further problem was the economics of installing a subsequent lining, necessitating setting up a headgear and hoisting arrangement approaching in size that required for conventional drill and blast sinking and lining. Because of the economics, German contractors opted for a phased transition from drill and blast to the full face, rodless, out of the solid shaft mole, by starting off with a down-the-hole shaft boring machine -without a drillstring-but using a pilot hole to get rid of the muck.

Jan 1, 1982

By W. C. Maurer

A drilling-rate formula for roller-cone bits is derived from rock crater-ing mechanisms. This formula holds for "perfect cleaning", which is defined as the condition where all of the rock debris is removed between tooth impacts. Under these conditions, the drilling rate is directly proportional to the rotary speed and to the bit weight squared, and inversely proportional to the bit diameter squared and to the rock strength squared. Under imperfect cleaning conditions, such m those usually present in field drilling, re-grinding of the cuttings occurs under the bit, and the drilling rates fall below those for perfect cleaning. INTRODUCTION Drilling with a roller-cone bit consists of two fundamental operations: the formation of craters under the bit teeth, and the removal of the broken rock from the craters. To delineate the mechanisms of drilling, it is first necessary to understand the mechanisms involved in the formation of individual craters and then to relate these mechanisms to the over-all drilling operation. CRATER FORMATION When a bit tooth impacts a rock, the rock is elastically deformed until the crushing strength of the rock is exceeded, at which time a wedge of crushed rock is formed below the tooth1, 2 as shown in Fig. 1. As additional force is applied to the tooth, the crushed material is compressed and exerts high lateral forces on the solid material surrounding the crushed wedge.- en these forces become sufficiently high, fractures are initiated below the tooth and propagate to the free surface of the rock. The trajectories of these fractures intersect the principal stresses at a constant angle,&apos;s4, as predicted by both Mohr&apos;s and Griffith&apos;s theories of failure. It has been experimentally 3,1 .6 that the volume of an impact crater, V,, varies as where Ec is the total energy imparted to the rock during the formation of the crater, and E is the threshold energy required before cratering is initiated (Fig. 2). Indexing experiments&apos; have shown that, when there is a second free face to which the crater can fracture, a larger volume of material is removed and that above a transition zone there is a linear relationship between V, and Eo., which has a slope equal to the relationship for the single crater. The beneficial effect of indexing tends to offset the threshold energy requirement to a large extent, as shown in Fig. 2, and the relationship is closely represented by Rate-of-loading effects have been used to explain the nonproportional relationships usually obtained between drilling rate and rotary speed. The data4, in Fig. 3 show that the Ec/Vc parameter remained constant for craters produced with impact velocities ranging from 10 to 8,000 ft/sec. Since no rate-of-loading effects were detected over this wide range of loading rates, it seems improbable that there would be detectable rate-of-loading effects present in the small changes in impact velocities produced by varying the rotary speed. The writer has shown (Fig. 4) that, for craters produced in rock by impacting spheres, where X is the depth of penetration. Simon2 has shown that this relationship also holds for craters produced in rock by penetrating wedge-shaped chisels. The energy used in crater formation is

By William F. Boericke

THE Christy Creek clay mine of the General Refractories Co., in the Olive Hill District, ranks with the most important producers in the north-eastern Kentucky fire clay field, both from production of high-grade flint clay, as well as unusual extent and thickness of deposit. The property is but one of a number controlled by the General Refractories Co. in this district, but is given first consideration among them by the operating staff. Its total area is over 5000 acres, owned in fee or by lease of clay mining rights, and although mining operations have been in progress for over five years, scarcely one per cent of the ground has been worked as yet. The mine is near the southwestern end of the clay field, which extends in a belt 12 miles wide and 55 miles long, in a southwestern direction from Kenova, W. Va., which, as the name indicates, is at the junction of Kentucky, Ohio, and West Virginia. Olive Hill, where there is one of the l1Ulin plants of the General Refractories Co., and to which the clay from the mine is shipped, is about 26 miles distant from the mine. Connection is made with the Chesapeake & Ohio R. R. by a spur line . 8 miles in length, owned and operated by the company.

Jan 1, 1929

By R. H. Sweetser

THERE are SO many conditions affecting blast-furnace results that it is hard to get satisfactory comparative data on the working of two furnaces, and much more difficult to get comparable results from the use of two entirely different fuels. The several advantages of charcoal over coke as a blast-furnace fuel have doubtless been apparent to many managers; but probably the conditions for comparative tests of the two were never so favorable as during the year 1905, at the works of the Algoma Steel Co., Sault Ste. Marie, Ontario, which comprise two modern blast-furnaces, well equipped, though of comparatively small capacity. For four consecutive months one of these furnaces was run with charcoal and the other with coke. Then the charcoal-furnace ran for a while with part charcoal and part coke; and, still later, with coke only. The ore-mixtures were about the same in the two furnaces, and, of course, all the climatic conditions were identical. The dimensions of the two furnaces, and their work with the different fuels, are given in Table I.

Jan 5, 1908