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By John T. Crawford

Much has been written of late regarding the flotation of coal as a means of reclaiming the valuable portion of the fines encountered in nearly all methods of coal preparation. Whether the process be wet or dry, whether the dust be sent through the entire process or removed prior to the cleaning by some such method as screening or aspiration, one of the final products consists of a considerable proportion of — 48-mesh material. The percentage of this —48-mesh material depends upon the mining method, transportation, storage, and method of preparation. The tonnage of this dust material ranges from 1 to 10 per cent of the total output of the cleaning plants. Four methods of disposal are available for this coal: 1. Disposal with the refuse from the plant. 2. Addition, after being reclaimed from the wash water, or dust collectors, without cleaning to the larger sizes. 3. Separate disposal as a pulverized fuel. 4. Cleaning so that addition to the larger sizes benefits the product as a whole. It is also true that flotation in one form or another is the only known way to effectively clean this material. While some processes do a reasonable cleaning job down to and including +l00 mesh, by far the greater majority stop at or short of the + 48-mesh size. Much of the literature on coal flotation passes by the fred pulp density and size by merely stating that the feed varies from 20 to 30 per cent solids and that the size range is from — 1/4 in. or — 1/8 in. to 0. Small wonder then that some operators with entirely different conditions, and not knowing the machine capacity, throw up their hands in despair when flotation is mentioned. Too much parallel is drawn between flotation methods as applied to ores and as applied to coal, to the detriment of the latter. In the flotation of metalliferous ores the feed pulp density ranges from 20 to 30 per cent solids, depending upon the amount of mineral to be floated and upon the size analysis. Thus where 2 to 5 per cent of the

Jan 1, 1936

By John T. Crawford

Much has been written of late regarding the flotation of coal as a means of reclaiming the valuable portion of the fines encountered in nearly all methods of coal preparation. Whether the process be wet or dry, whether the dust be sent through the entire process or removed prior to the cleaning by some such method as screening or aspiration, one of the final products consists of a considerable proportion of — 48-mesh material. The percentage of this —48-mesh material depends upon the mining method, transportation, storage, and method of preparation. The tonnage of this dust material ranges from 1 to 10 per cent of the total output of the cleaning plants. Four methods of disposal are available for this coal: 1. Disposal with the refuse from the plant. 2. Addition, after being reclaimed from the wash water, or dust collectors, without cleaning to the larger sizes. 3. Separate disposal as a pulverized fuel. 4. Cleaning so that addition to the larger sizes benefits the product as a whole. It is also true that flotation in one form or another is the only known way to effectively clean this material. While some processes do a reasonable cleaning job down to and including +l00 mesh, by far the greater majority stop at or short of the + 48-mesh size. Much of the literature on coal flotation passes by the fred pulp density and size by merely stating that the feed varies from 20 to 30 per cent solids and that the size range is from — 1/4 in. or — 1/8 in. to 0. Small wonder then that some operators with entirely different conditions, and not knowing the machine capacity, throw up their hands in despair when flotation is mentioned. Too much parallel is drawn between flotation methods as applied to ores and as applied to coal, to the detriment of the latter. In the flotation of metalliferous ores the feed pulp density ranges from 20 to 30 per cent solids, depending upon the amount of mineral to be floated and upon the size analysis. Thus where 2 to 5 per cent of the

Jan 1, 1936

By C. H. Stein

The Pennsylvania Legislature, on March 13, 1837, passed an act authorizing the Lehigh Coal & Navigation Co. to construct a railway to connect the North Branch Division of the Pennsylvania Canal with the slack water navigation of the Lehigh River; the navigation company accepted this act on May 10, 1837. Edwin A. Douglas, chief engineer for the navigation company, located the route for this railway from White Haven to Wilkes-Barre, which was called the Lehigh & Susquehanna Railroad. The object was to transport coal from the Wyoming Valley, of Pennsylvania, to the Atlantic seaboard. This railway was considered a marvelous piece of engineering, as it included a tunnel nearly 55 mi. long, and a series of inclines, or planes, up which the cars laden with coal were raised from the vicinity of Ashley to the summit of the mountain near Solomon's Gap, whence they were taken to White Haven by rail. There the coal was transferred to canal boats. In June, 1862, floods almost completely destroyed the upper division of the canal, so on March 4, 1863, the Pennsylvania Legislature passed an act prohibiting its restoration, but granted a charter for the construction of a railway from White Haven to Mauch Chunk to connect with the line from White Haven to Wilkes-Barre, previously built. On March 16, 1864, a supplementary act was passed extending the Lehigh & Susquehanna Railroad to Easton, Pa. This completed a direct all-rail route from Wilkes-Barre to the New York harbor (by way of the Central Railroad of New Jersey). Up to this time, all freight and passenger traffic passed over the planes, but in 1866 and 1867 there was constructed what was known as the "back track" down the mountain side between Solomon's Gap and Ashley over which all the passenger traffic and empty cars thereafter passed; it is now known as part of the main line. For the handling of the heavy coal traffic, however, this would have involved a continuous climb from Ashley, 170.27 mi., from New York, to Solomon's Gap, 157.8 mi. from New York, or a distance of 12.47 mi., and a total ascent of 1013.75 ft., composed of grades 95.61, 83, and 51.4 ft. per mi., respectively, and

Jan 1, 1922

By C. A. Hartnagel, D. H. Newland

Detailed production statistics listing the operations according to producing sands, districts, etc., are not available for New York State and are scarcely obtainable without extraordinary effort and expense, hardly warranted by the current rate of developments. The oil industry is practically stabilized, so far as exploration for new sources of supply are concerned; no new pools have been tapped in the last 25 or 30 years. There is little prospect of extending the boundary of the oil fields beyond their present limits, though possibly some additional reserves may be found below the present productive sands. Natural gas, on the other hand, is not restricted to such confined areas. Drilling for gas has been fairly active, under the stimulation of some important finds in formations that previously were overlooked. The year's field operations will be covered somewhat fully. Oil production, which has been steadily advancing for a number of years, showed a further moderate gain in 1936. The output for 11 months (figures for December are not yet available) was 4,217,000 bbl., indicating a total for the year of around 4,600,000 bbl. The increase was about 10 per cent, a rate maintained for some time. Prices were more satisfactory than in the preceding five years, with an advance to $2.57 a barrel before the year's close. New York State yield is all paraffin-base, in the same class with high-grade Pennsylvania oil. Most of the output is now obtained by the water drive, which has given a new lease of life to the fields and promises to bring the total to about 200,000,000 bbl. before the known acreage has been exhausted by this method. Before its introduction the annual production was under 1,000,000 bbl. a year. The fields cover an aggrcgate area of 55,000 acres, with some 17,000 wells in operation. Development Exploring for natural gas in New York in 1936 outside of the old established pools continued actively, as in the previous year. There was an increase in total initial production. Operations were conducted in nine counties of the central and western parts of the state, and 43 wells

Jan 1, 1937

By L. E. Porter, H. P. Hassel

The state of California produced 310,-996,696 bbl. of oil and about 415,832,000 M cu. ft. of gas in 1944. Such oil production represented 18.5 per cent of the nation's production, as compared with 19 per cent in 1943 ,The tremendous demand placed upon the California fields is indicated by the monthly production in 1944, as follows: Month Barrels per Day Jan................... 809,262 Feb.................. 820,712 Mar.................. 824,592 .4pr ................. 829,973 May.................. 832,068 June.................. 846,038 July.................. 849,231 Aug.................. 858,932 Sept.................. 885,823 Oct................... 877,667 Nov.................. 876,613 Dec.................. 883,847 In all, 2100 wells in proved areas and 269 wildcats were drilled in California during the year 1944. At the end of the year, approximately 270 wells were drilling in the state, as compared with 156 wells at the same time in the preceding year. Nine new fields or new zone discoveries were made during the year, which, according to best estimates, will contribute some 77,000,000 bbl. of new reserves; in addition, the extensions of old fields will increase the new reserves about 165,000,000 bbl., or a combined total, through dis- coveries of new fields, zones, and extensions, of 242,000,000 bbl., offsetting to the extent of about 78 per cent the total oil produced during the year. Geographically, the reserves from these new field and zone discoveries and extensions are distributed as follows: Per Cent San Joaquin Valley........ 72 Coast Area............... 24 Los Angeles Basin Fields. . . 4 100 The state's reserves of crude oil at the end of the year have been estimated to be distributed as follows: The outstanding development project for the year was that of the United States Navy in the Naval Petroleum Reserve No. I, Elk Hills field. During the year, 115 wells were completed in this field, and at the close of the year an average of 19 strings of tools were drilling. This program is being jointly conducted by the Standard Oil Co. of California with the Navy and is designed to increase the production of Naval Petroleum Reserves from approxi-

Jan 1, 1945

By Edward Steidle

Those who have followed the rapid progress in rock-dusting bituminous-coal mines must admit that opinion crystallized during the year just closed in support of this preventive of coal-dust explosions. Up to 1924 only one mining company in America practiced rock-dusting on a large scale, but during 1924 forty-two companies had adopted this safety measure, and the "boom" is expected to extend over the next two or three years. Alabama is represented by 1 company, Colorado by 3, Illinois by 4, Kentucky by 1, New Mexico by 4, Pennsylvania by 20, Utah by 8, and Wyoming by 1. A total of about 500 miles of entry has been dusted. Illinois leads with 275 miles of dusted entries and Pennsylvania holds second place with about 175 miles. In each case there is a sense of great security, not only on the part of the operator but by the miner as well. It is also interesting to note that during 1924 the compensation rating or inspection bureaus in twelve of our coal-producing states revised their rating schedules so that a credit would be given for rock-dusting. This credit ranges from 10 to 20 cents per $100 payroll. Recent drastic federal and state regulations approved by mine operators made rock-dusting compulsory in bituminous coal mines in Utah, beginning July 1, 1924. Legislative action favorable to rock-dusting may be expected in other states. Colorado has already proposed a change in its laws, and Pennsylvania, Maryland, Washington, and West Virginia are giving the matter consideration. If the rock-dusting problem is handled properly and proves to be entirely successful, no mining man of any standing will delay applying it in his mine. State legislation regarding this matter may or may not solve the problem. At present, the mining companies that are rock-dusting of their own choice are doing it carefully and thoroughly. In Utah, some are doing it conscientiously, trying to get what benefits can be obtained while others are reported to be doing it only to keep within state regulations. In any event, if we are to have new legislation, it is essential that the various states know what they are about and then, if possible, work for uniform regulations.

Jan 1, 1925

By A. Lawley, H. W. Schadler

TWINNING has long been recognized as a possible mode of deformation in crystalline solids and has been studied in a wide variety of crystals.&apos; Recently, deformation markings which have the topographical characteristics of twins have been seen in tungsten" and the body-centered cubic alloys of tungsten and molybdenum with rhenium.3-7 These markings have been identified as twins because: 1) they are visible after polishing and etching;2,3 2) their formation is accompanied by an audible click and a drop in load;4 and 3) the habit plane is {112}, the plane which has been identified as the twinning plane for several body-centered cubic metals.&apos; However, there has been no published evidence for the specific determination of the twinning direction in rhenium alloys of tungsten and molybdenum, because usually the deformation twins were not broad enough to be detected with a conventional X-ray beam. Consequently, [111] is inferred as the twinning direction since the latter must lie in a plane of symmetry perpendicular to the twinning plane (112). As Sims and Jaffee3 and Lawley and Maddin4 have shown, the addition of rhenium to molybdenum and tungsten increases the ductility of the alloy over that of the pure metal, but twinning is much more profuse in these alloys than in the pure elements, particularly at Mo-35 pct Re*and W-30 pct Re. Since twinning is often associated with brittle fracture in iron8 and the silicon-irons,9 direct experimental proof that the deformation markings in tungsten and molybdenum alloys of rhenium are twins seemed to be needed. Using the Laue back-reflection technique, the deformation markings in a Mo-35 pct Re alloy are shown to be deformation twins. The twins have a (112) habit plane and the twinning direction is <111>. The Mo-35 pct Re specimen used was taken from a 0.060-in. zone-refined single crystal deformed 6 pct in tension at 78°K. The Laue back-reflection technique, with a film to specimen distance of 2 cm, and white radiation from a copper target were used.

Jan 1, 1962

By G. I. Madden, L. H. Van Vlack

WhEN basic refractories which contain peri-clase, MgO, are used in steel furnaces, they are altered by iron oxide to form magnesiowiistite, (Mg,Fe)O, and an iron-rich magnesioferrite, (Mg,Fe)Fe2O4. Silica, which is usually present, produces a silicate liquid that surrounds the solid phases. The individual crystalline grains of the above two solid phases grow at elevated temperatures. Previous papersL" showed that, if either periclase or magnesiowiistite is present as the only solid phase, its grain size increases in proportion to temperature and in proportion to the square root of time; the growth rate is reduced slightly if the amount of silicate liquid is increased. Similar results have been observed for iron-rich magnesioferrite.3 To determine the effect of the presence of two solid phases on the rate of grain growth, micro-structural studies have been made on samples containing both magnesiowustite and magnesioferrite. The experimental procedure was the same as that followed in the previous studies of this type.&apos;12 Briefly, this included determination of compositions from the available phase relationships4 that would produce the two phases of interest at the firing temperature and using reagent-grade raw materials to prepare mixtures of these compositions. The mixtures were pressed into pellet-size samples and then fired at the desired time and temperature. The samples were water-quenched to preserve the microstructure present at the firing temperature. Standard reflected-light metallographic techniques were used for sample examination and sample analysis. Grain sizes were determined by measuring the mean diameter of a number of randomly selected grains in a two-dimensional microsection. A large cross section of the sample was covered to prevent biased data resulting from such effects as localized differences in the radius of curvature of the crystal grains. Examination indicates that grain growth is restricted when more than one solid phase is present. This is illustrated in Fig. 1 for: (a) magnesiowustite plus a silicate liquid; (b) magnesioferrite plus a silicate liquid; and (c) magnesiowiistite and magnesioferrite plus a silicate liquid. Fig. 2 plots the

Jan 1, 1964

By C. H. Mathewson

UNFORTUNATE evasions of metals specifications recently brought to public attention through news items and editorials have caused executives of at least two great corporations to set up defensive procedures and to re¬affirm the patriotic purpose and conscientious effort of their plants in this battle of production. Whatever the exact conditions affecting these cases, in which the public interest has been safe-guarded by vigorous action, my reaction to such offenses in general is a feeling that they arise from the pressure and limitations that afflict our overextended war effort and that inevitably there will be more of them, doubtless more and more expertly conceived and less and less subject to detection as simple faked, or substituted analyses or tests. To the ladies, who doubtless are already beginning to ac¬quire a sense of painful resignation to a dry technical talk, may I say that the cookbook is full of specifications which customarily are varied within rather wide limits but it is generally recognized that the proof of the pudding is in the eating without regard to its exact composition or some scientific evaluation of its texture.

Jan 1, 1943

By Porter W. Shimer

The device here described has been found useful in sampling foundry-iron, and there is no reason why it should not be equally useful in sampling other metals, which are not too hard to be drilled with a breast-drill. It consists, as shown in Figs. 1 and 2, of an ordinary plumber&apos;s gas T (without bead) of about 1/2-inch in internal diameter, over the leg of which is snugly fitted a small tin cup, about If inches deep, to catch the drillings which are made by use of a breast-drill with a 1/4-inch twist bit, working through the arms of the T. The forward arm of the T is filed down to a blunt edge, over which is stretched a short piece of rubber tubing projecting a little beyond the end. The bottom of the inside of the front arm of the ? is filed down to a slope, to facilitate the passage of drillings into the tin cup. The rear arm of the T is fitted with a tin ring, projecting about 1/4-inch, for the purpose of holding in place a stiff spiral brass spring, which serves to press the sampler firmly against the metal to be drilled. The other end of the spring is supported against a shoulder of the breastdrill. The projecting rubber prevents the loss of any drillings, and also the falling of sand from above into the sampler. With ordinary care there need be no risk of getting particles of rubber into the sample. It is, of course, not possible to get a fair average sample of drillings, representing a pile of foundry-iron, by taking drillings from any one or two pieces, because of the frequently mixed character of the iron. This observation applies more particularly to the sampling of iron as it is broken and piled at the foundry, than to sampling at the works, where, if the furnace is working regularly, a few pigs may fairly represent a whole cast. By the use of this device, it is possible for a sampler to pass

Jan 1, 1901

By C. Y. Cheng

THE purpose of this note is to show that the dislocation mechanism controlling the creep of Fe-4 pet Si alloy1 may equally account for the behavior of a-Fe2 over the same temperature range. A recent study of the creep behavior of a-Fe has been made by Ishida et al.2 Their results were in agreement with the mechanism of nonconservative motion of jogged screw dislocations over the entire range of temperatures investigated, from 480° to 1045°K. However, the observed creep rate was much higher than that which could be accounted for by the above model unless the authors assumed that pipe diffusion became significant below 775°K. More recently, Cheng et al.1 inves-gated the creep of Fe-4 pet Si alloy over about the same temperature range and found that the results could be correlated fairly well with a dislocation climb mechanism as represented by Eq. [l]. Proceeding as shown previously,1 if creep is controlled by a single climb mechanism, at intermediate stresses, the shear creep rate, y, can be represented by where n and K are material constants {K is independent of the temperature but dependent on the structure, that is, K decreases during the primary stage and reaches a steady-state value over the secondary stage of creep), t is the applied shear stress, G is the shear modulus, gc is the free energy of activation for creep (usually equal to that for self-diffusion, gd), and RT has the usual meaning. By solving for y, the apparent activation energy for creep, Q which can be determined experimentally, is given by where hc is the enthalpy of activation for creep which usually is also equal to that for self-diffusion, h^. By using Eq. [2], the data in Fig. 2 of Ref. 2 can be re-plotted as shown in Fig. 1. By comparing the two figures, it is seen that the slight increase in the apparent activation for creep over Region II, as noted in Fig. 2 of Ref. 2, is due to the influence of temperature on the shear modulus rather than of the stress on the

Jan 1, 1969

By George Hallerman, R. J. Gray

In the customary method of studying age hardening, the process of aging is interrupted by cooling the specimen and measuring its room-temperature hardness. However, the aging process may be conveniently measured while it is occurring at an elevated temperature by a series of hardness measurements at suitable time intervals. The latter method gives data void of complications that might result from processes occurring during the quench, and the data may be more indicative of the alloy&apos;s properties at elevated temperatures. The elevated-temperature hardness tester1 used in the present study employed a dead-weight loading system with a synthetic sapphire Vickers-type in-denter mounted in a molybdenum holder. The tests were performed in a vacuum of l0 to 10 torr. Hardness impressions were made by applying a 2-kg load for 15 sec. Although the hardness impressions made in the tester at elevated temperatures were measured at room temperature, the correction for thermal contraction during cooling was insignificant and therefore neglected. The two methods of obtaining age-hardening data were compared using eighteen samples of Haynes alloy No. 25* which had been annealed for 2 hr at 2250°F (1232°C). The test surfaces of the 1 by 1 by 1/4-in. samples were ground and then polished on vibratory polishers2 prior to testing. Three groups of five samples were aged at 1400°, 1650°, and 1800°F (760°, 899", and 982°C) for 4, 16, 32, 64, and 100 hr and then water-quenched. Room-temperature hardness as a function of aging time for these samples is shown in the broken lines of Fig. 1. Only one sample for each of the three temperatures was necessary to determine the hardness of the alloy at the aging temperatures with the elevated hardness tester. The specimens were heated isothermally and hardness impressions were made after various time intervals up to 100 hr. The results of these determinations were represented by solid lines in Fig. 1. Comparison of the two sets of curves shows, as expected, the hardness values determined at the aging temperatures are lower than the room-temperature hardness of the quenched specimens and the hot-hardness values reflect a strong tempera-

Jan 1, 1964

By A. S. Eve

IN the spring of 1930, the question was raised as to the possibility of estimating the depth to which the pyrrhotite-nickel deposit at the Falconbridge mine extended in the earth. This body is 7500 ft. long, it dips at 80° to the north, and it is fairly straight with slight sinuosities. Dr. F. W. Lee suggested that it might be possible to solve the problem by measuring the vertical magnetic intensity, with a good magnetometer, both near the surface of the earth and also at a level a &apos;few feet higher in the air. The writer regarded this suggestion with some scepticism, but it was decided to put the scheme to a prac¬ tical test. In July, 1930, Dr. Lee, Dr. Keys and I &apos; the writer, with Mr. Joyce and Mr. Belloc as assist¬ ants, made a firm, light wooden platform (Fig. 1) E measuring 4 by 4 ft. on the top, which was 10 ft. above ground. This platform could be moved easily a few yards by four men, and carried across I country in a wagon or by motor truck. Some iron nails were used in its construction, but we were not able to detect their influence on the - Askania magnetometer. Nevertheless, in future work, the use of iron should be completely avoided. Readings were taken along a traverse at right L angles to the orebody, at several stations, both on FIG. 1.-VERTICAL the ground and on the platform. The excess read- ASKANIA MAGNETOM¬ ETER ings, over the dike, as compared with those about 10 ETER ON A FT. HIGH. PLATFORM 800 ft. away, are shown in Table 1. The largest reading, 6.8, is open to suspicion and should be rejected. The depth of the top of the orebody is known by electrical survey, and by actual drilling, to be about 112 to 123 ft., and the covering is sandy glacial drift. Similar readings were taken over a long straight magnetic diabase dike. These also are shown in Table 1. The magnetic vertical intensity is consistently less at the higher level, which is l0 ft. above the lower level.

Jan 1, 1931

By T. R. Braithwaite

The brief comments made in this chapter deal mainly with shaft sinking, but don&apos;t assume that freezing techniques apply only to shaft sinking. They can be used on tunnel work, on such things as liquid gas storage in sediment, and I think you will probably see other applications for it as we go along. The method was first developed in 1883 and was used to sink shallow shafts through water-bearing overburden in Germany and in England. One of the early shafts was in Germany and was sunk by freezing to 310-ft depth. The project required 17 months for drilling, roughly 5 1/2 months for freezing, and another 7 months for sinking, giving an average progress of 10 ½ fpm. Fortunately there have been numerous improvements and new developments in this technique through the years. With equipment and techniques available today, we can take a 20-ft-diam shaft down through 2000 ft of water-saturated overburden and the time schedule runs roughly 4 months for site preparation, freeze-hole drilling, installing the casings, and setting up the plant; 4 ½ months for freezing; and 6 months for excavating and installing permanent linings. This gives a progress rate of about 140 fpm, on average. I might qualify that a little in that the special linings that are required for some formations can affect the time. If it&apos;s straightforward sinking where one can use a heavy concrete lining, up to 630-640 fpm may be achieved in a shaft; but with special cast-iron tubbing or types like that, 5 fpd is pretty good progress, so the lining does affect the time. The freezing method is used mainly for sinking shafts through unstable water-bearing formations which cannot be grouted either effectively or economically and which are too deep to enclose in pile casing. Even with overburden, grouting methods available today can handle roughly 100 lb water pressure; in the range of, say, 100 to 150 ft from the surface, it&apos;s very often best to grout 40 or 50 ft. You normally drive a pile casing and go through it. Below that range, freezing is very often the economical

Jan 1, 1970

By L. O. Howard

The practice of using acid-lined converters at the plant of the Old Dominion, Copper Mining & Smelting Co. was discontinued early in January, 1913, and replaced by basic-lined converters. The new equipment consists of one electrically operated 12-ft. upright converter stand and two basic-lined shells for same. These shells are provided with twenty-four 1 1/4-in. tuyeres. The mouths are 5 ft. in diameter. Owing to the fact that there is no bedding system or other large storage capacity between the mine and the smelter, the furnaces must handle as fast as produced ore and concentrates, which may vary considerably from day to day, both in tonnage and copper content. Consequently the converter is often crowded to its utmost capacity by the frequent fluctuations in the daily matte supply. From the foregoing, it will be seen that in no way can the &apos;converter be favored by running light charges or by transferring to it from other stands in operation. Neither can it regularly be allowed to cool between blows, and only occasionally can it be run at as low a temperature as might be wished. No copper precipitate or other foreign material high in copper is available to help out the tonnage. Repairs and alterations or the removal of defective tuyeres must wait until such time as a slight falling off in.the matte supply allows the converting operations to ease up. These variable conditions all combine to interfere with the systematic operating of the converter. The first 12-ft. shell was blown in on Jan. 6, 1913, and ran until June 27, 1913, when it was removed to patch the brick. A record of operations covering this period follows: Shell No. 1. First Campaign, Jan. 6,to June 27, 1913 Commenced blowing............................Jan. 6, 1913. Taken off stand for patching.....................June 27, 1913. Total blowing time..........................2,689 hr. 40 min. Total number of, blows made.............................. 664 Average time of blow............................ 4 hr. 3 min. Tons matte charged.................................. 18,157 Average copper content of matte................ 43.9 per cent. Average time blowing one ton. copper................ 22.5 min. Total copper made............................... 7,250 tons

Jan 1, 1915

By Berke L. Thompson

INTRODUCTION Purpose of Wall System There were 2379.88 m (meters) of tied back wall along Interstate 77 in the Capitol Complex area in Charleston, West Virginia. The main purposes of this wall system was for aesthetics, reduction of excavation, saving the forested slopes, and control of sloughage of the back slopes onto the Interstate. Failure At 1 PM on Thanksgiving Day, 1982, 3 wall panels fell from the insitu position onto the interstate in a matter of seconds. See Fig. 1. The total length of wall that fell was 29.25m. The height of the three panels varied from 10.06m to 10.36~1 and each individual panel had a length along the highway of 9.75m. These were abutting panels that were approximately in the middle of a section of wall 165.81m long. The panels separated from the remainder of the wall at 2 expansion joints that occur at the ends of every 3rd panel. The reinforced portland cement concrete panels were designed on a 76.2mm (millimeter) batter. See Fig. 2 for Wall Dimensions. The footer for these panels was lm wide, was keyed into sandstone bedrock 0.30m and was approximately 1.52m below the portland cement concrete pavement and shoulder. Thirteen of the 17 panels in this section of wall were tied to the sandstone bedrock behind the wall by either 25.4mm or 34.93mm hollow rock bolts. Two panels on each end of the wall were cantilever type walls. The panels that failed had 25.4m hollow rock bolts. A portion of each panel adjacent to both sides of the failure was pulled out from the rock face and appeared to be in imminent danger of also failing. All of the 21 rock bolts in the panels that fell were separated either at the backface of the wall, in the extension from the wall to the rock face or at the rock face. None of the rock bolts had been pulled out of the bond length in the sandstone bedrock. After considering the condition of the wall and

Jan 1, 1984

By David J. Mack, Dennis R. O’Boyle

The mechanical properties of three polycrystalline intermediale Phases that have the y bvass structure were measured in compression between 400° and 900°K. At the lower testing temperatures— termed Region I— no plaslic deformalion occurred prior to brittle fracture. A1 higher temperatuves— lermed Region III—the three phases deformted plaslically for all strain rules from 1.5 to 36.0 x 10-4 sec-1. The homologous temperature for the inilialion of plaslic flow increased in the same order as tile atomic size difference between atoms in each plmse (0.50Tmp for y AgZn, 0.62Tmp for y CuZn, and 0.76 Tmp for y CuCd). The increased flow resistance oj. y CuCd above 0.50Tmp is explaitzed in terms of the more random arrangement of copper and cadmiwn atoms on the laltice sites of&apos; the y bvass slrilctlire. Plaslic deformation of the three intevmiediate phases in the smooth plaslic flow region appears to be controlled by oacancy-induced dislocation climb. RESULTS are reported in this paper on the mechanical properties of three binary intermediate phases having the complex cubic y brass structure (D82, 143m) containing fifty-two atoms in the unit cell. The structure of the prototype y brass phase, y CuZn, was first analyzed by Bradley and Thewlis&apos; based on the X-ray work of Westgren and Phragmen.2 The y brass structure can be visualized as an arrangement of twenty-seven body centered cubes (3 by 3 by 3) into a unit cell containing fifty-four atoms. From this cell the center atom and the four corner atoms are removed and the remaining fifty-two atoms are slightly displaced, resulting in the y brass structure. Approximately thirty-five binary intermediate phases that have an electron to atom ratio of 21:13 have been reported to form the y brass structure. The three y brass intermediate phases selected for this study are formed from elements in groups IB and IIB of the first and second long period. Each compound exists over a composition range that includes the stoi-chiometric composition A5BB (Cu5Zn8, Ag5Zn8, Cu5Cd8). y CuZn (a0 = 9.944A) has solubility limits extending from 57 to 67 at. pct Zn at 500°C apd melts peritecti-cally at 834°C; y AgZn (a0 = 9.326A) has a solubility range from 58 to 63 at. pct Za and melts peritectically at 661°C. y CuCd (a, = 9.596A) melts congruently at 563°C and has a maximum range of solubility extending from 58 to 64 at. pct Cd. In order to minimize coring and microporosity during solidification, the composi- tion of the compounds that form peritectically (y CuZn and y AgZn) was chosen so that the temperature difference between the liquidus and the solidus is only a few degrees centigrade. In selecting y brass compounds for this study, phases were avoided that contained a transition element or that had an electron to atom ratio greater than 1.70. Hume-Rothery et a1.3 has observed that, when the electron to atom ratio of y brass intermediate phases exceeds 1.70, atoms begin to drop out of the unit cell to maintain a constant electron to unit cell ratio. EXPERIMENTAL PROCEDURE Each of the three intermediate phases was prepared from elements having a purity >99.999 pct. Only trace quantities of impurities (<0.0001 pct) were detected by spectrographic analysis of the four elements. A master alloy of each compound weighing approximately 130 g was prepared by melting the elements in evacuated sealed quartz tubes heated in an electric furnace. Specimens having the desired diameter for the mechanical property measurements were obtained by remelting the master alloy in a Vycor tube under argon and drawing specimens from the melt into 4-mm-diam quartz tubes. To obtain a smooth surface on specimens drawn from the melt, the inside of the quartz tube was coated with a thin layer of graphite formed by thermally decompositing acetone. All specimens were imbedded in sealing wax and cut to the proper length with a diamond-impregnated cut-off wheel. The individual specimens (4 mm diam by 8 mm long) were then sealed in evacuated Pyrex tubes and homogenized at 500°C for 72 hr. Chemical composition, variations over the length of the rod (10 cm long) were generally less than 0.2 pct and the analyzed composition differed from the desired composition by less than 0.5 pct.4 Metallographic evaluation of the specimens after homogenization showed no evidence of a second phase in any of the compounds. The average grain size of the specimens after homogenization was 1 to 2 mm. All mechanical property measurements were carried out in an argon atmosphere using a compression apparatus described previously.5 Prior to applying the load to the specimen, the test apparatus was evacuated, back-filled with argon, and heated to the test temperature for 20 min to establish thermal equilibrium. EXPERIMENTAL RESULTS Mechanical properties measured as a function of temperature and strain rate for each of the three y brass phases were similar. From room temperature to the ductile-to-brittle transition temperature, only elastic deformation was observed prior to fracture.

Jan 1, 1968

By Beverley S. Randolph

While our coal-mining practice, in regard to hauling on main roads, has advanced very rapidly in recent years by means of compressed air, electricity and ropes, that of gathering from rooms or working-places has remained almost stationary. Few large mines are without some method of mechanical haulage on main roads, but the gathering from working-places is still done almost entirely by means of animals or men. An effort to improve this has been made recently by the Consolidation Coal Company, under the direction of the writer, at its mines in the Georges Creek region. The seam worked is the Pittsburg bed, known locally as the " Big Vein." It is from 8 to 12 feet thick. Immediately overlying the coal there is from 5 to 6 feet of " rashings," consisting of thin alternating beds of shale and coal, which disintegrate rapidly on exposure to the air, and making a very treacherous roof. Rooms are driven from 12 to 15 ft. wide, with a single track close to one side. A line of posts is placed just far enough from this side to leave a clear space for the mine-car with a driver at the brake. This brings the post not far from the middle of the cross-bar, and provides more effectual support than if placed at the end. The tracks in these rooms are usually of 4 in. x 4 in. oak scantling, though of late years, owing to the advance in the price of lumber and the reduction in steel rails, more of the latter are being used. This track is laid by the miner. His pay for the work is included in the price per ton for mining. It is therefore, as a rule, unskillfully and often carelessly laid, and cannot be relied upon to carry safely any weight materially greater than the loaded mine-car. Compressed air was already in use on the main roads, driving motors weighing 30,000 pounds each, which were used to. bring the mine-cars that had been assembled from the rooms

Jan 1, 1904

By William R. Webster

The unusual interest taken in the papers on steel at the New York (1922) meeting showed that the time is ripe for the renewal of the general discussion of the physics of steel, on the same lines that wete used formerly and produced such good and lasting results. The curves shown in Figs. 1 to 4 and the appendix were added to my paper1 after the February meeting, but, unfortunately, Figs. 1 and 3 were transposed; the curves on page 252 are for bessemer steel and belong on page 254; while those on page 254 are for open-hearth steel and belong on page 252. These errors are evident at once on referring to the data in Tables 1 and 2, pages 227 and 228. The curves for each 0.05 per cent. of carbon and manganese in basic open-hearth steel are given in Figs. 1 and 2; these lines for the lower carbons and manganese are comparatively uniform. Figs. 3 and 4 give similar lines for carbon and manganese in bessemer steel. The large break in the carbon lines from 0.55 per cent. to 0.60 per cent. is caused by much higher silicon in the steels of 0.60 per cent, carbon and over that are used for springs, etc. The estimated ultimate strength of pure iron, and the increase in strength by carbon, phosphorus, and manganese, with the different views regarding the same, were given and diseussed. All of this can be considered as an introduction to the practical use that is now being made of such data in selecting the heats of steel best suited to give the required physical tests called for in the finished rolled material. The chemical composition of the steel, of course, tells only a part of the story, as many other factors that have an influence on the physical properties of the steel must be considered, and controlled as far as is possible. Some of these have already been referred to, but there are other important ones to be considered and discussed.

Jan 1, 1923

By J. E. Schaefer

A summary of oil and gas development in Ohio for 1936 is necessarily brief and, in some details incomplete, owing to the lack of a central state agency for the collection of data and statistics on well drilling and production. The information used in this report was compiled from the records of the several companies operating in the state. Drilling activity in Ohio increased during 1936 as compared to the two preceding years. During the year 1157 wells were completed, as compared to 1022 wells in 1935 and 1055 wells in 1934. However, in spite of this increase in drilling activity, both oil and gas production declined for the year. Activity by Fields Trenton Sand Field.—The Trenton sand field is located in 15 counties of northwestern Ohio, trending from Toledo on the north, south and west through Findlay into Indiana. About 40 years ago the Trenton field was one of the most important oil-producing fields in the United States, but it has declined to a position of minor importance. During the year 1936 there were 52 completions in this field—29 gas wells, 19 oil wells and 4 dry holes. Clinton Sand Field.—The Clinton sand is by far the most important gas-producing field in the state. Production is located in small pools separated by nonproducing areas, on a band of varying width reaching from Cleveland on the north to Lawrence County on the Ohio River on the south. During 1936 a small pool of patchy character was developed in Butler township, Knox County. Seventeen wells were completed in the Clinton sand, of which 13 were gas wells and 4 were dry holes. The average initial production secured per well drilled, including dry holes, was 1,288,000 cu. ft. The average depth of the Clinton sand was 3000 feet. A gas well was completed in lot 9 of Franklin township, Tuscarawas County. Production was found in the first Clinton sand at a depth of 4354 ft. The initial open flow after shot was 828,000 cu. ft. and the rock

Jan 1, 1937