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By Thomas H. Clagett

ThE Pocahontas coal field comprises the area in Tazewell County, Va. and Mcrcer and McDowell counties, W. Va., in which Nos. 3 and 4, Pocahontas seams of bituminous coal are mined. It is a mountainous region with ridges rising 300 to 1000 ft. (91 to 305 111.) above the very narrow valleys and, except for some poorly kept farms along the tops of the ridges and the area occupied by the mining and coking plants and villages, is wooded throughout. First shipments were made from Pocahontas in 1883. There are now about 100 mines, 95 per cent. of which are drift mines. At the end of 1920, approximately 298,000,000 gross tons of coal had been produced and 35,000 acres of coal mined. No. 3 seam is from 4 to 10 ft. (1.2 to 3 m.) thick, but the change is gradual and the thickness fairly uniform in any particular locality. It has a streak of bone, about 2 in. (5 cm.) thick, to which the coal adheres on both sides. In parts of the field this bone is replaced by hard slate, usually from 2 to 4 in. thick. The seam has a slate or fireclay bottom and a draw-slate top. No. 4 seam is from 65 to 80 ft. (19.8 to 24 m.) above No. 3 seam, and of minable section only in the western part of the field. It is from 3 to 8 ft. thick and has two streaks of bone, cach from 1 to 2 in. thick. It has a fireclay or slate bottom and is overlaid by from 4 to 24 in. of a laminated coal and slate, locally termed "black rash," which is separated from the coal seam by about 3 in. of slate. This "black rash" averages about 2.5 per cent. ash, but contains some clean coal. The coal in both seams is quite soft. Cleavage sufficieiltly well defined to exert an influence on the mining is rare. Overlying strata arc slates, shales, and sandstones, the sandstones predominating, extending from 30 to 1000 ft. above No. 3 seam. The strata directly above the draw slate of No. 3 seam, or "black rash" of No. 4 seam, and up to the first sandstone are usually slates or shales, about 10 ft. thick over the.

Jan 1, 1923

By R. P. Carreker

ONE of a series, this report is concerned with the experimental documentation of the deformation behavior of pure metals over a wide range of temperature. Previous reports in the series describe the creep behavior of platinum1 and aluminum2 and the tensile behavior of copper,&apos; silver,& and molybdenum.5 Subsequent reports will describe the creep behavior of copper and silver and the tensile behavior of platinum, thus providing extensive creep and tensile data on four face-centered-cubic metals tested under comparable conditions. Material Two lots of high-purity aluminum were used in this investigation. One lot, designated AC, was obtained direct from the Aluminum Co. of America as high-purity aluminum in notch-bar form. A typical analysis of this type of aluminum is 0.006 pct Si, 0.015 pct Cu, 0.006 pct Fe, and 99.975 pct Al. Lot BS was obtained from J. I. Hoffman of the Bureau of Standards as a portion of the Aluminum A described in his report of the redetermination of the atomic weight of aluminum.6 Lot BS originated with the Aluminum Co. of America also, and was received in the standard notch-bar form. A detailed chemical and spectroscopic analysis of lot BS has been published.6 Principa1 impurities were 0.006 pct Si, 0.003 pct Fe, 0.002 pct Cu, and >99.987 pct Al. Samples were processed identically by room-temperature swaging and drawing to 0.030 in. diam wires from approximately ½ in. diam rods that were machined from the notch-bar pigs. Samples were cut from the cold-drawn wire, placed in a grooved graphite block in groups of thirteen, and annealed in air for 1 hr at each of several temperatures, with the results shown in Table I. The two lots of aluminum responded quite differently to annealing, the behavior of aluminum AC suggesting that its grain growth was controlled by impurities. The recrystallization texture of aluminum of similar purity comparably treated has been reported as predominantly <111> fiber with some <100> fiber present.7 Testing Procedure The testing procedure and method of analyzing the data have been described previously.2,3 Annealed specimens of 5 in. gage length were tested in an Instron tensile testing machine at a constant head motion corresponding to a strain rate of 0.04 min-&apos; at a number of temperatures. Instantaneous tenfold rate changes were employed in some tests to deter-

Jan 1, 1958

By M. H. Richman, L. M. Markowitz

MUCH of the available information on which the heat treatments of a particular steel are based is found in the isothermal-transformation diagram. Such is the case despite the fact that the usual thermal treatments involve continuous transformations, i.e., transformations which occur on cooling, rather than isothermal transformations. This rather unusual practice is continued because of the abundance of published isothermal-transformation diagrams1, 2 which are relatively easy to determine experimentally as compared to continuous-transformation diagrams.3, 4 However, this can easily lead to the production of improper micro-structures and may not allow full advantage to be taken of the inherent hardenability of a steel. A mathematical method of estimating the continuous-transformation diagram from the available isothermal data was developed by Grange and Kiefer;5 but, due to the mathematical approximations required in the calculations, the computation of continuous-transformation diagrams has been limited to individual cases and no large-scale compilation of these diagrams is available. With the present availability of high-speed digital computers, the problem of time-consuming approximations may be minimized and continuous-transformation diagrams may be constructed with the insertion of isothermal data into a basic program. It is the purpose of this report to announce the availability of such a program and compare the computed and experimental continuous-transformation diagrams for two different steels. Let us first review the method of Grange and Kiefer5 and then compare the results of the machine calculations with the experimentally determined diagrams. In order to apply this method, one requires the isothermal-transformation diagram and a set of curves representing cooling at constant rates. A section of an isothermal-transformation diagram is shown in Fig. 1 together with a cooling curve of M0F per sec. The cooling curve is constructed to start from the A1 (eutectoid) temperature if the transformation product is the eutectoid microconstituent (i.e., pearlite or bainite), from the A3 temperature if the product is primary ferrite, or from the 4, if it is proeutec-toid cementite. The cooling curve in Fig. 1 intersects the curve representing the start of isothermal transformation at a point A&apos; corresponding to a temperature TX and a time fX. An arbitrary lower temperature To (on the

Jan 1, 1968

By Robert G. Woolery, S. Ramachandran, James A. Weber, Donald J. Hansen

Union Carbide began looking seriously at heap leaching in 1971. At that time some 1.6 million tons of mineral averaging 0.40 kg U3O8 /t) (0.80 lb U3O8) were stockpiled at various sites around the Gas Hills, Wyoming, uranium facility. Several alternatives for economically exploiting these reserves were considered and heap leaching seemed the most attractive. At the time, Western Nuclear Inc. was operating a heap leaching project nearby in which the low-grade ore was leached with sulfuric acid, the product liquor treated in an on site solvent extraction circuit, and the stripped product liquor re-acidified and recycled. The operators were most generous in allowing us to visit their site and in answering our questions. They were also helpful in discussing the problems associated with heap leaching. Their experience confirmed our belief that it was possible to operate a heap leach profitably on material too low grade for conventional milling. Based largely on what we had learned from Western Nuclear&apos;s experience, we undertook a program involving a laboratory study to determine the characteristics of the mineral stockpiles, as related to heap leaching, and to conduct a pilot heap leach adjoining our Gas Hill mill. Basically the pilot heap construction and leach procedure were similar to those described in the literature l, 2 with certain modifications to permit detailed monitoring of the operation. In May 1972, the decision was made to construct a test heap. The UCC plant at Gas Hills, Wyoming was selected as the most appropriate site for the following reasons: (1) The necessary equipment for heap construction was readily available. (2) Stockpiles of mineral from the various mining sites were on hand: (3) Construction of the heap could be made on an abandoned tailings pile minimizing problems associated with environmental approval. (4) Proximity of the mill would allow product streams to be absorbed into existing circuits. (5) Necessary utilities were available. (6) Tailings water from the existing circuit could be used for leaching. A program for the pilot heap leach was developed with the following primary objectives. (1) To confirm the data obtained in bench-scale tests particularly on uranium recovery, acid usage, and ore permeability. (2) To evaluate plastic liners used for base preparation. (3) To investigate heap configuration variables. (4) To determine heap leach amenability of the different mineral sources. (5) To establish the economics of heap leaching low-grade ores.

Jan 1, 1979

By completing a double-track railroad cut-off, involving a large volume of cut and fill, in addition to a half-mile bridge spanning an important French river, engineers in the zone of the Services of Supply have opened up a route by means of which hours of valuable time are being saved in transporting men and materials to combat areas. Its most valuable feature is that it provides a means for by-passing a certain city where several, French railroad lines converge and where, consequently, congestion of traffic has been responsible for delayed train movements. To avoid a double-track turnout, which would have involved a cross-over cutting one of the French tracks, the two track connections of the cut-off are taken off from opposite sides of the French right-of-way. This involved running one of the new lines of the cut-off out from the main line track on a fill to a point where it could be brought back and over the French track on a steel girder bridge. Beyond this bridge the two tracks of the cut-off converge and straighten out in the standard type of double-track construction. The new cut-off is about 5 3/4 mi. (9 km.) long and has required 160,000 cu. yd. of cut (122,329 cu. m.) and 414,000 cu. yd. (316,527 cu. m.) of fill. In the biggest continuous fill a-yardage of 180,000 has been placed; the road-bed in this section-being-about 40-ft. (12 m.) above the original ground surface. The bridge portion of the cut-off, 2190 ft. (667.5 in.) long, is made up of ninety-nine 14-ft. (4.3 m.) timber spans and sixteen 50-ft. (15.3 m.) steel girder spans. This is the longest bridge which engineers of the American Expeditionary Forces have built in France. The timber spans are supported by pile trestle bents, while the longer steel girders rest on clusters of 43 piles. Two pile drivers of the steam-hammer type worked from each -end of the bridge toward the center of the river, while a drop hammer worked westward from the shore end. In this structure nearly two thousand piles 50 ft. or so in length had to be driven.

Jan 12, 1918

The Library belonging to the Institute is now administered as part of that of the United Engineering Society, the members of the Library Committee representing the Institute on the Library Board of the United Engineering Society. The main collection now consists of 115,900 books, 3680 pamphlets, 5811 maps, and 4402 searches, making a total of 129,793, a net gain of 7701 within the year. There is in addition a lending collection of 1004 volumes and one of duplicates amounting to 11,310. The library grows steadily both by gift and purchase. The cataloging is now substantially current covering 124,100 items represented by 362,145 cards. There are many rare and notable books, especially those relating to early engineering. It is truly a great treasure house of knowledge. While no severe effort is made to record all visitors the books showed 23,470 to have used the Library in 1928. In addition 5102 telephone requests for information were answered, 202 searches and 196 translations made and 34,964 photo prints furnished. The service department is maintained by small fees which cover the cost. The total amount appropriated for operating expense for the year was $46,300, of which $8,240.02 was furnished by the Institute, $5,399.45 being derived as income from the Douglas Library Fund and the remainder from current receipts. Within the year a slight change was made in the method of meeting the expenses of the Library, a flat charge of $5,000 per year being made against each of the four Founder Societies and this being supplemented by a surcharge based upon membership. The effect of this has been for the present to decrease slightly the amount that the Institute would otherwise need to pay. Placing in one collection all the books of the four societies greatly increases the number of books accessible to the members of each and opening the Library to the public is an important factor in keeping the tax exemption privilege enjoyed for the building. The Director and Staff of the Library make every effort not only to accommodate those who come in person to consult the collection but members resident elsewhere who communicate by letter or wire. The facilities of the Library are truly available to all.

Jan 1, 1929

The State Mining Experiment Station, Missouri School of Mines and Metallurgy, Rolla, Mo Charles H. Fulton, Director of the School of Mines Martin H. Thornberry, Director of the Mining Experiment Station. From the State Mining Experiment Station are issued all the general and technical bulletins of the School of Mines The latest bulletin containing a complete list of publications will be sent upon application to the Director of the Station Residents of the state may obtain bulletins as issued without charge, or may ask to be placed on the mailing list. Nonresidents should write to the Director for further information concerning specific bulletins. From the long list of bulletins the following have been care¬fully selected as being most likely to interest the geologist and the mining engineer. Vol 8, No. 1 (1916), Bibliography of concentrating ores by flotation, by J Cunningham, vol 1, No 3 (1912), Some apparatus and methods for demonstrating rock drilling and the loading of drill holes in tunneling, by L E Young, vol 2, No 1 (1915), Comparative tests of piston drill bits, by C R Forbes and L M Cummings; vol 2, No 3 (1916), Cupellation losses in assaying, by H. T Mann and C. Y Clayton; vol 4, No 1 (1917), The hydrometallurgy and electrolytic precipitation of zinc, by F D. James; vol 4, No 2 (1917), The effect of additional agents in flotation, pt 1, by M. H. Thornberry and H T. Mann, vol 5, No 1 (1919), The carbonization of Missouri cannel coals, by H L. Dunlap, K. K. Kershner and V. X Smiley. Vol 5, No. 4 (1921), Coal mining methods in Missouri, by W W Weigel, vol 6, No 1 (1921), The problem of the St Peter sand¬stone, by C L Dake, 50 cents, vol 6, No. 2 (1922), Bibliography on the electrothermic metallurgy of zinc, by B M. O&apos;Harra, vol 6, No. 4 (1923), Experiments on the distillation of zinc from complex zinc-lead-silver ores, by B M. O&apos;Harra and E S. Wheeler, vol 7, No 3 (1924), Underground loading devices, by C. E. van Barneveld, 81; vol 7, No. 4 (1924), Briquetting of zinc ores, by B. M. O&apos;Harra; vol. 8, No 2 (1925), A treatise of Missouri clays, by M. H. Thornberry.

Jan 1, 1933

By A. M. Nicholas

WheN attempting to lift mill pulp containing a considerable percentage of wolframite, in an ordinary bucket elevator, difficulty was encountered from the tendency of the tungsten minerals to settle, on account of their high specific gravity. Lifts not exceeding 20 ft. (6 m.) could be accomplished without particular difficulty, because there was insufficient time for consolidation of the heavy minerals in the bottom of the bucket; but with higher lifts, the heavy tungsten minerals would settle so solidly that when the bucket turned over at the top of the elevator the contents could be only partially dumped. In this case it was necessary to lift the material a distance of 34 ft. (10m.); hence some means had to be adopted to prevent settling, at least to such an extent that the material could be dumped from the buckets. To reduce the carrying time to a sufficient extent would have required a belt speed of 475 ft. (144.7 m.) per minute, which was not feasible. It was decided to try, by vibrating the belt about midway of its span, to keep the material sufficiently agitated during elevation to prevent settling. Underneath the rising side of the belt was placed a 12-in. (30.48-cm.) idler pulley, 18 ft. (5.5 m.) up in the total lift of 34 ft. The idler pulley was set eccentric by % in. (19 mm.) and it was hoped that the vibration thus imparted to the belt, running at a speed of 280 ft. (85 m.) per minute, would prevent settling in the buckets. It was found that the pressure of the belt on the idler, due to the small inclination of the elevator, was not sufficient to overcome sliding on the idler, and the agitation was insufficient to accomplish the desired purpose. The idler was then connected to the drive shaft of the elevator through an intermediate pulley, gearing down to a speed of 20 r.p.m., so that a positive vibration was imparted to the belt; this imparted enough motion to the belt to prevent settling and the contents of buckets dumped satisfactorily. Protecting Elevator Belts When elevating, from the crusher, a rock which was very sharp and abrasive, considerable difficulty arose from cutting of the belt under-

Jan 1, 1918

By Frank A. Wardlaw

MY subject covers the effect that recent metallurgy has had on operations at the Utah-Delaware mine. This mine is the old Highland Boy mine of Bingham Canyon, Utah, one that has now been in operation for about 30 years. It was first operated by Messrs. Weir and Newhouse as a gold mine, under direction of the Highland Boy Gold Mining Co. The ore was oxidized gold ore and it was treated on the property in a cyanidation plant. During the process of development, sulfide ore was found assaying about 5 per cent copper. It proved to be the discovery of what is yet the largest sulfide copper orebody in the State. The first shipment was made in December, 1896. A smelter was built in Salt Lake Valley to treat this ore and not long after the property was sold to the Utah Consolidated Mining Co. From 1900 to 1924, the mine was a large producer of direct-smelting ore which averaged for many years 4.5 per cent. copper, then later dropped to 2 per cent. copper. During the years 1912 to 1920, direct-smelting lead ore was also mined. In 1918&apos;and 1919 ore developments at depth showed a decided increase in silica content and a porphyry sill so closely associated with the ore that it was impossible to mine the ore clean and it was found necessary to build a copper flotation Eoncentrator at Tooele. The mill had a capacity of 1000 tons a day and made a recovery of 90 per cent., and would have been responsible for a continuance of profitable operations had it not been for a slump in the copper market and the necessary high cost of mining ore on the lower levels. RESULTS OF SELECTIVE FLOTATION In 1924 the mine was on the point of a shutdown which had all the earmarks of being permanent; it was being operated at a heavy loss and preparations were made to pull the pumps and salvage the machinery. The plant at Tooele was remodelled for selective flotation by the Inter- national Smelting Co. in order to treat custom ores, and incidentally to clean up a possible 10,000 tons of lead-zinc ore which it was thought possible to get in the Utah-Delaware before shutting the mine down. Results were so encouraging that an effort was made to develop more lead-

Jan 1, 1928

By John Henry Harden

THIS shaft, located in the northern anthracite coal-field about 2300 feet southwest from the court-house at Wilkes-Barre, in the County of Luzerne, Pa., is the property of the Lehigh & Wilkes¬Barre Coal Company. The sinking and cribbing to the solid rock (see isometrical view on Plate IX) was conducted by Mr. W. W. Kenrick. Dimensions of the earthwork to the foundation of the cribbing, 51&apos; 1/2" long, 17&apos; 4" wide, 31&apos; deep. Total amount of excavation in cubic yards, 998.5. Dimensions within the permanent cribbing, 45&apos; 4" long, 11&apos; 16" wide, 31&apos; deep, divided into six parts by buttons 8" x 12", each division, 7&apos; x 11’ 16". Permanent cribbing 12" x 12"; temporary cribbing 8" x 10"; backing, 3" thick ; clay puddle to ex¬clude surface water, 12" thick; conductors, 8" x 5"; lining, 1" thick from top to bottom of the shaft. In the latter part of the year 1871, Mr. H. Murray & Son contracted with the company for the continuation of the work at the rate of four hundred dollars per yard in depth, to a point 100 feet below the Hillman seam of coal ; for the succeeding 100 feet the price to be increased twenty-five dollars per yard, and so on until the bottom rock of the Baltimore seam of coal was found. The size of the shaft within the cribbing timbers is 45 feet 4 inches by 11 feet 6 inches, having six divisions of equal dimensions, 7 feet by 11 feet 6 inches, giving room for what is known as the" standard diamond" car, whose capacity when level full is 100 cubic feet, but when loaded by the miner 4 inches above the top contains 112 cubic feet.* Two divisions of this shaft are intended for hoisting coal from the Hillman seam, two from the Baltimore seam, one for ventilation, and one for pumps. The works were planned for two pairs of direct-acting, high-pres¬sure engines, twenty-four boilers, and for an estimated output of 2500 tons of coal per day from the two seams. * Size of the "standard diamond" car, 8&apos; 10” x 4&apos; 3" x 2’ 8” = 100.1.4; loaded 4 inches above the top, 8&apos; 10” x 4’ 3” x 3’ “0"= 112.7.6.

Jan 1, 1877

By K. P. Anderko

INVESTIGATIONS of the constitution of the system Mg-Ba were first made by Grube and Dietrich1 between 0 and 47 atomic pct Ba and then by Klemm and Dincke1acker" ver the whole composition range. Klemm and Dinckelacker corroborated intermediate phases of the approximate formulas Mg9Ba and Mg4Ba, and showed a third compound to have the formula Mg2Ba, rather than Mg5Ba, as suggested by Grube and Dietrich. Mg2Ba had been previously recognized by Hellner and Laves3 s a MgZn2 (C14) type structure. The main discrepancies between the results of Grube and Dietrich and those of Klemm and Dinckelacker exist, however, in the liquidus temperatures and solidification equilibria of alloys containing more than about 10 atomic pct Ba. The liquidus temperatures measured by Grube and Dietrich, and the outline of the liquidus of Klemm and Dinckelacker are plotted in Fig. 1 on an enlarged scale. Although there are strong reasons to prefer the results of Klemm and Dinckelacker, it was considered worthwhile to remeasure liquidus temperatures of some alloys in the controversial composition range. The alloys, made up from sublimed magnesium and 99.5 wt pct pure barium, were melted in an Armco-iron crucible in a dynamic argon atmosphere. Thermal analyses were taken using a Speedomax auto- matic recorder. Careful stirring during the solidification process was required in order to avoid segregation. After thermal analysis some of the cast alloys were chemically analyzed. The intended and analyzed compositions of the ten alloys prepared are listed in Table I. Since standard barium determinations cannot differentiate between barium present as element and as oxide, both the barium and magnesium contents of the alloys were analyzed. The fact that the sum of barium and magnesium was close to 100 pct in each case was taken as evidence that no excessive oxidation had occurred. The liquidus temperatures measured are listed in Table I and plotted in Fig. 1. As can be seen, they are in good agreement with the liquidus found by Klemm and Dinckelacker. This leaves no doubt that the thermal results of Grube and Dietrich above 10 atomic pct Ba are incorrect. This work was performed under the sponsorship of the Pitman-Dunn Laboratories of Frankford Arsenal, Philadelphia, Contract No. DA-11-022-ORD-1975. Sponsorship and permission to publish by this organization are gratefully acknowledged.

Jan 1, 1958

By Robert M. Betts

The Cornucopia district is situated in northeastern Oregon 25 miles from the Snake River branch of the Oregon Short Line R.R. and 75 miles northeast of Baker City, in the Wallowa Mountains. The formation is principally granodiorite, which has intruded much older rocks locally known as greenstone, and which are probably altered schists. In the upper horizons of the veins, greenstone is found as inclusions in the granitic mass. Both of these formations have been cut by numerous basalt, aplite, and porphyry dikes. Veins of the district are the result of faulting and the consequent crushing of the wall rock. The ore has been deposited irregularly along these faults and formed a lenticular type of vein. The ore varies in width from a few inches to 20 ft.; the average stoping width is about 5 ft. As the basaltic dikes are subsequent to the ore deposition, they are at times a hindrance in development and mining operations; on the other hand, a dike which has followed a vein as a line of least resistance often serves as a good mall in place of the crushed, rather loose granite. The veins generally strike northeast-southwest and dip about 35" to the west. Orebodies are from 100 to 300 ft. in length and about the same in height. The lenses are irregular; inclusions of wall rock are common in the ore and are a factor in reducing the grade. The ores are quartz, carrying some free gold near the surface, but the metals for the most part are in iron sulfides, sometimes associated with chalcopyrite. The gold and silver are in a mechanical combination with the iron, which requires find grinding to liberate them. The ratio of silver to gold is 5 to 1, by weight. Development The Cornucopia Mines Co. owns, and operated until 1920, two mines and two mills of 100 tons capacity each. The sustained high costs and labor shortage, however, necessitated the abandoning of the Union mine, the oldest in the district. Two crosscut tunnels were then started, first, to cut the Last Chance vein at a much lower horizon and, second, to prospect two parallel veins. The ore from the Last Chance vein will go to the old Union mill and the ore from the new veins to the Last Chance mill.

Jan 1, 1925

By William Hutton Blauvelt

The type of gas producer in which the ashes are fluxed and run off as slag was among the very earliest made. Ebelmen built the first one in 1840 at Audincourt, France, only a year after the installation of the first gas producer of which we have record. Charcoal was used as the fuel, with blast-furnace slag and clay as flux. My interest in this type of producer began when as a boy I saw the fluxing producer at Chester, N. J., which was invented by W. J. Taylor, and described by him in 1881.&apos; This producer was invented independently by Mr. Taylor to meet the difficulties he experienced in making producer gas for roasting sulphurous iron ores. It will be of interest to review briefly his description of his producer. It was built like a small blast furnace, having a hearth 24 in. in diameter and 24 in. high. The bosh angle was 25&apos; from the vertical; the diameter at the bosh 4 ft. and at the top 3 ft.; the total height 12 ft. The producer contained one water-cooled tuyere 12 in. above the bottom, with a 1.5-in. nozzle. Depth of the coal above the tuyere, 6 ft. Mr. Taylor mixed the coal with from 30 to 40 per cent. of basic blast-furnace slag to flux the ash. Occasionally some limestone was also used, but never limestone alone, as the use of the furnace cinder gave a larger volume of slag and made it easier to maintain a proper fluidity. The producer was blown with a small Weimer blowing engine, delivering 300 ft. of air per minute, at from 1 to 1.5 lb. pressure. The slag was tapped every 2 hr. and was black and glassy in appearance. Broken or egg size anthracite was used. The fine sizes of anthracite could not be made to work. Mr. Taylor expressed the belief that bituminous coal could be used if not too fine, but no experiments were made with this coal. Runs were made of four weeks&apos; duration, with no stops or changes. Mr. Taylor reported the following advantages: 1. Excellent gas, very uniform in quality. The high fuel bed permitted no air to pass unconsumed, and the gas was almost entirely free from carbonic acid.

Jan 1, 1914

By T. H. Alden

In several recent studies of fatigue fracture in single-phase metals, tests have been made using the condition of essentially constant plastic strain amplitude.1-4 This procedure departs from the more common practice of holding the amplitude of stress constant. From these studies have emerged two important generalizations about the behavior of metals during cyclic straining. First, measurements of the stress amplitude, required to impose the fixed plastic strain amplitude, have shown that the rate of cyclic strain hardening in annealed metals is rapid in the early cycles but decays quickly to a very low value.1"5 This behavior is displayed particularly for a range of strain amplitudes below about 2 pct and at low and high temperature. Frequently, the hardening rate becomes so low that one can determine a "saturation" stress. This stress increases with strain amplitude. In addition, an experimental relation between fracture life and plastic strain amplitude was obtained by Coffin and Taternelli,&apos; namely N 1/2 ep = constant. For several metal:;, this equation has been shown to hold down to a strain amplitude of 0.1 pct. Whether it also holds in the "low amplitude" range (lives of more than about 105 cycles) has not been established. The importance of the saturation stress level (a value derived from strain hardening measurements) to the question of fracture life (the usual quantity of interest of fatigue testing) lies in the observation that cracks start to form when the cyclic hardening rate is essentially zero, and not before.4"5 This point suggests immediately a means of relating tests at constant stress and constant strain amplitudes. When making comparison, the stress in a constant stress test is to be identified with the saturation stress in a constant strain test. If these ideas are correct, then it should be possible, using the relation between fracture life and strain amplitude and experimental curves of saturation stress versus strain amplitude, to predict the fracture life as a function of stress, i.e. predict the S -N curve, for low stress fatigue. Saturation stress amplitude-plastic strain amplitude data, obtained in tension-compression, exist only for a few metals and for relatively high amplitudes of strain. Comparable S -N curves (push-pull loading) were known by the author only for OFHC copper. Values for this metal taken from Ref. 2 are shown in Fig. 1. The strain amplitudes, the lowest

Jan 1, 1962

By Paul Gordon, A. S. Iyer

Apreliminary investigation has been carried out on the annealing behavior of explosively deformed copper as compared to that of conventionally deformed copper, using hardness, stored energy, and the sharpening of X-ray spots on Debye rings as the measures of annealing. Some metallographic observations were also made. Two purities of copper were studied. The first was tough pitch copper obtained as a small piece from the guard ring of a sample explosively deformed to approximately 500 kbars.* The second consisted of mentedand deformed these samples quite badly so that only a few small fragments were available for the annealing experiments. As a result, little in the way of quantitatively significant data was gleaned from the measurements. However, several important qualitative differences between the annealing of explosively deformed and conventionally deformed copper have been observed. It is these effects which are described in this report. Fig. 1 presents the results of hardness and stored energy measurements during the annealing of the high-purity copper after explosive and after conventional deformation. The data for the conventionally deformed copper were taken from the work of Gordon;&apos; in this copper it is evident that most of the stored energy is released concommittantly with the hardness drop. As shown in orddon&apos;s&apos; investigation, these effect:: are definitely associated with recrystal-lization. In the case of the explosively deformed copper, on the other hand, a large fraction of the stored energy is released before any appreciable hardness change takes place. At 160°C about 40 pct, and at 170°C about. 80 pct, of the stored energy appears to have been released prior to the onset of softening, i.e., undoubtedly prior to recrystallization. In view of the facts that the hardness and energy measurements were made on separate samples and that there is reason to suspect that the intensity of the explosive deformation from sample to sample may not have been uniform, a check run was carried out in which both energq release and hardness changes were measured cnn the same sample. The results of this run, shown in Fig. 2, confirm the conclusion that much of the stored energy is released before the hardness begins to change appreciably. In an effort to obtain some idea of the activation energy associated with the early release of energy, a single callorimetric sample was annealed in two steps, first at 149.5"C and then at 159.5"C, Fig. 3.

Jan 1, 1962

By G. P. Walters, W. C. Fuller

THE production of thin films of beryllium became essential early in the irradiation program at Harwell in order to study the fundamentals of irradiation damage in this metal. In common with other metals, electron-microacope specimens of beryllium can be produced by different techniques. The techniques used to produce thin-film specimens varied according to the starting material. Most of the starting material was -0.060 in. thick. Large-area specimens (1 in. long by 0.5 in. wide) were chemically thinned to -0.005 in. thick in a solution containing 85 pct orthophosphoric acid, 10 pct chromium trioxide, and 5 pct sulfuric acid at 70°C prior to electropolishing. Small-area specimens, Fig. 1, had to be mechanically thinned prior to electropolishing because chemical thinning considerably reduces the area as well as the thickness. Mechanical thinning tends to produce a worked layer about 0.002 in. deep and hence this method cannot be used to thin bulk specimens below about 0.005 in. thick. Mechanical thinning is carried out by mounting the specimen on perspex then carefully polishing on wet grinding papers. Polishing is continued until about half the initial thickness is removed; the reverse side is then polished in a similar manner. After the bulk specimens have been mechanically or chemically thinned to -0.005 in. they are elec-tropolished in the solution shown in able I at 50" to 80°C. This solution should be prepared by adding the chromium trioxide to the orthophosphoric acid and then the glycerol last. The glycerol should be added slowly because a vigorous effervescence can occur. Electropolishing is carried out by holding the specimen in stainless-steel tweezers and gently agitating in the solution. Masking or "window" techniques are not required for beryllium, good microscope specimens being obtained from the edges of elec-tropolished foils. A stainless-steel beaker should be used, acting as the cathode and solution container. A closed circuit voltage of 30 v is used for polishing. Too low a voltage produces a matt surface and pitting occurs when it is too high. The solution temperature is important. If it is too low, the solution is viscous and bad polishing results and, if too high, etching of the specimen occurs. After electropolishing, the foils are thoroughly

Jan 1, 1963

By B. E. Lindsly

The withdrawal of gas and/or oil in appreciable quantity from a natural oil reservoir causes the pressure within the reservoir to diminish, and if the oil is completely saturated with gas, this dissolved gas begins to evolve from solution. The liberation of gas from solution in oil causes the oil to change in several ways, the principal changes being: (1) reduction in volume or "shrinkage"; (2) increase in specific gravity (decrease in degrees A.P.I.); (3) decrease of energy in the oil; (4) increased viscosity; (5) increased surface tension. As the writer&apos;s investigations have been concerned more with the first three items than with changes in viscosity and surface tension the following discussion relates particularly to the volumetric shrinkage of the oil, to the increased specific gravity, and to the decreased energy in the oil, each caused by the liberalion of dissolved gas. One of the most surprising facts that have developed in the study of high-pressure well-head and "bottom-hole" samples is the great reduction in volume that occurs in some oils when the naturally dissolved gas is liberated. High-pressure samples taken from well heads in the Oklahoma City, Kettleman Hills and Ventura, Calif., fields decreased in volume from 11.2 to 40.5 per cent when the dissolved gas was liberated in reducing the pressure on the sample to atmospheric. The amount of shrinkage an oil undergoes when gas is liberated depends upon the quantity of gas liberated and also upon its composition. A given volume of the heavier gases, propane and butane for example, will occupy more space in the oil than an equal volume of methane. The reason for this probably is that the molecules of the heavier gases are larger than those of the lighter gases. When gases are not in solution, and particularly when they are at low pressures, the distance between the gas molecules is so much greater than the dimensions of the molecules themselves that the mere size of the molecules has little effect upon the space occupied by the gas, Avogadro&apos;s hypothesis, which in effect states that all gases under the same conditions have the same number of molecules per unit volume, relates to this phenomenon. The fact that the heavier gases occupy more space than the lighter gases in the oil, and

Jan 1, 1934

By F. E. Bash

PREVIOUS to the war, this country depended on Europe for its supply of a number of alloys of great importance in the manufacture of electrical apparatus and equipment. When this source was cut off shortly after the beginning of hostilities and the small stocks in the United States were exhausted, either inferior substitutes had to be used or the manufacture of the apparatus suspended. Since electrical equipment was of primary importance in the prosecution of the war, the substitutes had to be used until the composition and processes of manufacture of the alloys could be determined. Included among these alloys was one generally known as manganin, which is composed chiefly of copper and manganese and which, beside having a low-temperature coefficient of resistance, has a low thermal electromotive force against copper. These qualities make it valuable for the manufacture of electrical instruments and machines in which it is used in the shape of wire, rod, and strip. As a result of these circumstances, an investigation was started to deter mine how manganin could be made on a commercial basis so that a product of uniform quality and the necessary electrical characteristics could be produced. The object was to procure an alloy with temperature coefficient of resistance not greater than T 1 X 10-5 per ° C. and a thermal e.m.f. against copper no greater than that of the imported material. It is characteristic of manganin that the temperature-resistance curve is not a straight line but has a peak, which varies in position with different samples; it is most desirable, however, to have it at a temperature of 25° C. if possible. The coefficients of manganin from all records available ranged from a large negative to a large positive value. The composition as reported by different authorities is given in Table 1. Prof. Matthew A. Hunter, of Rensselaer Polytechnic Institute, made up a number of small melts of manganin in a graphite crucible and the one that he reported as having the best coefficient gave the following analysis: Copper, 83.6 per cent.; nickel 2.5 per cent.; manganese,-13.6 per cent.; iron, 0.3 per cent. An analysis of a sample of German manganin gave the following: Copper, 82.6 per cent.; nickel, 3.8 per cent.; manganese, 12.8 per cent.; iron 0.7 per cent.

Jan 9, 1919

By Terje Preber

Undisturbed sampling of shale has always been difficult. As pert of a comprehensive subsurface investigation for a nuclear power plant in northwestern Illinois, the Maquoketa shale, which contains seams of argillaceous dolomite, was sampled using different coring techniques. Since the Maquoketa shale is extremely fissile, careful sampling preservation by both lateral and vertical confining was utilized. The samples were tested to determine both static and dynamic properties. A relationship between void ratio and compression index, swell index, and swell pressure ratio was established. Correlation was found between the dynamic properties and both undrained shear strength and confining stress. Based on the results of the sampling and testing programs, recommendations for sampling and sample preservation techniques are given. INTRODUCTION Undisturbed sampling and preservation of samples of weathered, fiasile shale is often a very difficult task. The often erratically varying degree of weathering with depth complicates sampling and makes reproducibility of testing results difficult. Therefore, the testing of weak, weathered rock very often yields a wide variety of test results with large data scatter. The reproducibility of teat results usually associated with water-deposited soils is seldom, if ever, obtained. The testing and sampling program described in this paper was part of a comprehensive subsurface investigation for a nuclear power plant in northwestern Illinois. The sampling, preservation and testing program of shale samples taken from the Maquoketa shale are described, and, in particular, the different sampling techniques and sample preservation procedures are evaluated with respect to their effectiveness. The samples obtained by these procedures were subjected to a comprehensive testing program, including both static and dynamic testing in order to provide input for soil-structure interaction analysis. GEOLOGY The Maquoketa Shale Group is of Ordovician age and is part of the Cincinnatian Series. The area where the shale samples were taken was in a broad sea far from sources of clastic sediments during most of Ordovician time. However, during Cincinnatian time when large deltas in the Appalachian region were being formed, clastic sediments from the east dominated the sedimentation ii. Illinois, resulting in formation of the shale and dolomite shales of the Maquoketa Shale Group (Willman et al., 1975). The geologic unit sampled consists of a greenish gray to gray silty shale containing numerous 1- to 12-inch (2.5- to 30.5-cm) thick interbeds of argillaceous dolomite. The dolomite comprises up to 40 percent of the profile studied. The shale

Jan 1, 1984

By Harold R. Bradford

EXPANSION initiated during and after the war has placed industrial plants in new areas and increased reduction and manufacturing facilities in communities already established. With added expansion interest has grown in possible area contamination from waste products. Evolved flue waste contains many chemical contaminants that may or may not cause concern; however, atmospheric pollution has always been a problem to industrial communities. For some time effects attributed to small quantities of compounds of fluorine in various waste gases have been studied by industrial, medical, agricultural, and associated interests.1&apos;2 Aside from an article by Churchill, Rowley, and Martin in 1948," little has been published on the subject of fluorine in coal except by foreign investigators. The first indication of fluorine compounds in coal was noted in England in 1934 by R. Lessing,4,5 who found that the presence of fluorine in Midland and West County coals was definitely established by etching of glass and by severe corrosion of porcelain fillings in a gas works scrubber. Etching of glass in a direct contact annealing kiln initiated research, and in 1944 four papers were published by H. E. Crossley,6-9 giving results of investigations of fluorine in coal of England by the Fuel Research Board, Department of Scientific and Industrial Research. These investigations studied methods of determining fluorine, the quantities present in various coal deposits, the geographical distribution of fluorine in British coals, occurrence of fluorine in minerals associated with coal, amounts of fluorine compared to other constituents of coal, the nature of fluorine compounds, and industrial significance of fluorine in coal. It was shown that British coals contain from 5 to 200 ppm of fluorine, which is probably present as fluorapatite [Ca10F2(PO4)6], often written as [3Ca3P2O8.CaF2]. The fluorapatite is intimately associated with shale in some coals and can often be separated from these coals along with the shales. H. E. Crossley8 also reported no correlation between the amounts of chlorine, sulfur, or ash and the amount of fluorine in different coals. He found, however, a relationship between the content of fluorine and phosporus in many British coals. Thus it should be possible to anticipate likely fluorine-rich coals from existing data giving phosphorus content. Behavior of the fluorine compounds of coal was compared with that of 12 fluoride-phosphate minerals in sink-float separations and solvent ex-

Jan 1, 1958