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By J. E. Lawrence

Copper-decorated lattice disorders in silicon have been analyzed by electron transmission microscopy, chemical etching, and P-N junction reverse current measurements before and following different gettering treatments. Lattice disorders acted as the nucleating sites for copper precipitation. Copper precipitates in silicon resist chemical etching, absorb electron beams more than silicon, and cause excess reverse currents. Totally gettered disorders will chemically etch rapidly, be transparent to electron beams at extinction contrast, and not contribute to excess reverse currents. Three different gettering procedures were performed to determine the effectiveness of different EXTRANEOUS metal contamination can alter the minority carrier lifetime and resistivity of bulk silicon as well as cause excessive reverse currents in semiconductor devices. Previous workers have reported such activity for gold,1,2 copper,'-* iron,' and rnanganese.2,5 Gettering is a method employed to render these impurities electrically inactive. However, some property of the junction-degrading metals prevents gettering procedures from being wholly reliable. The comparative effectiveness of gettering with different agents has been the topic of earlier publications.2,3,6-11 Two mechanisms for gettering metallic impurities have been proposed: the distribution coefficient mechanism by Thurmond and Logan7 and the compound formation mechanism by Goetzberger and Shockley.2 Both of these models evolved from data gettering mechanisms. First, phosphorus was allowed to enter a protective SiO2 film but remain away from the Si-SiO2 interface to test the effectiveness of the compound formation gettering mechanism. Second, phosphorus was allowed to enter the backside of silicon wafers to test the effectiveness of the lattice strain gettering mechanism. Finally, nickel was diffused into the backside of silicon wafers to test the effectiveness of the distribution coefficient gettering mechanism. In each case, copper precipitates dissolved as a result of the gettering treatment. Tightly packed precipitates will resist gettering more than loosely packed copper precipitates. acquired by the electrical analyses of gettered semiconductor materials. An electrical analysis, however, cannot identify the fate of the gettered metals. The radiotracer technique has recently been applied to determine the redistribution of metals following different gettering treatments.3,9,10 The spatial resolu tion of the radiotracer technique prevented the analysis of metal precipitates which contribute to the excessive reverse currents of P-N junctions. The gettering processes employed by the earlier workers, almost without exception, allowed the gettering agent to enter the silicon lattice. This procedure makes it quite difficult to differentiate between the distribution coefficient and the compound formation mechanisms. The metallographic analysis of gettered copper precipitates discussed in this work supplements the earlier work by obtaining direct evidence of silicon lattice defects decorated with copper before and following different gettering treatments. The effectiveness of a process which restricts the gettering agent, phosphorus, to a dielectric film and thus away from the silicon lattice is discussed in addition to the ef-

Jan 1, 1969

By P. R. Sperry, M. L. Holzworth, P. A. Beck

The basic work of Z. Jeffries 1,2,3 has long ago established the main features of grain growth in the presence of a dispersed second phase. Working with sintered specimens of initially fine grained tungsten, to which various amounts of thoria had been added, Jeffries found that grain growth was inhibited, that is, practically prevented, up to a certain annealing temperature. When this temperature, which increased with the thoria content, was exceeded, extremely large grains developed abruptly from the fine grained matrix. such a a temperature" was later found by Gross-mann~3,14 in certain types of steel, while ~~i~ showed15 that in other steels grain growth was gradual, without inhibition and abrupt coarsening. Later work demonstrated the connection between coarsening and aluminum additions. The viewpoint that the direct cause of inhibition and coarsening in aluminum killed steels is a fine dispersion of aluminum oxide21 appears to be held quite generally, although some doubts have been voiced even very recently, 20 In steels with Ti additions the titanium carbide phase is considered responsible for the inhibition effects found.20 The coarsening temperature increases with the titanium content. That coarsening may occur, as a result of, certain heat treatments, even in low carboil rimmed sheet steel, was shown by Samuels. He also detected coarsening in a steel ingot of similar composition, after giving it the same heat treatment as that used by him for the sheet material. In this instance the identity of the inhibiting phase has not been determined. However, Tangerding found23 that even the few small carbide particles, which occur in carbonyl iron, have a very marked inhibitive effect. As a result of an oxidizing anneal at 850°C, large grains begin to form at the surface of the carbonyl iron specimen, and gradually grow inward as the subsurface oxidation of the carbide particles progresses and eliminates the obstruction. The carbide particles Can be removed also by annealing in hydrogen, with similar results. But if the oxidizing anneal of 10 hr at 850°C in air is followed between in hydrogen for 75 hr at the Same temperature, grain growth is more hibition usually the whole specimen is transformed into a single crystal (approx. 1.5 cm X 5 cm). The reversed procedure, that is, hydrogen annealing followed by oxidizing annealing leads to smaller grains. Tangerding arrived at the logical conclusion that there must have been in his specimens some additional obstruction to grain growth hibition from a minor impurity, other than carbon. It seems likely that this obstruction was caused by an oxide formed during the oxidizing anneal and gradually reduced rtain the subsequent hydrogen-annealing. Other instances of grain growth inhibition of varying severity, associated with a dispersed second phase, have been described by several authors. Notable examples: lead with 0.06 pet Cu or Ni,24 cartridge brass with 0.12 pct Cr,25 or with up to 0.15

Jan 1, 1949

By Archibald Macmartin

THE Königin-Marien-Hütte is the only works in Germany where the Bessemer process is carried on by the direct method. The Besserner plant there, is arranged after the true English type, and the only resemblance to the Swedish mode of procedure is the dispensing with the use of spiegeleisen at the end of the "blow." In a new department of the establishment, started within three years, each of the converters is turned by means of a very neat and compact reversible engine, the steel shaft of which is an endless screw, which turns against the oblique cogs of a large wheel attached to the shaft of the vessel. An advantage which this arrangement possesses over the ordinary English hydraulic arrangement, is the fact that the endless screw suffices to turn the vessel, in either direction, any number of complete revolutions; while even the latest American improvements, so clearly explained to us by Mr. Holley at the opening session of this meeting, do not secure even one complete revolution without changing the angle of inclination of the hydraulic piston. In the endless screw arrangement, there being no limit to the working of the motor in either direction (no return-stroke necessary), the vessel can always, unless outside reasons demand the contrary, be turned around to any desired position by the shortest cut, whether backwards or forwards. Also, the diameter of the cog-wheel attached to the vessel can be made sufficiently great to avoid all unevenness of motion. In the old department the motive-power still continues to be hydraulic. If reports be true, the new department produces spiegelized steel, for the manufacture of all-steel rails. But the old department is still,' as from the beginning, devoted to the production by the direct method of steel for steel-headed rails. The most remarkable fact connected with this direct steel is the ease with which it welds to the iron' of the rail-packets, although no borax or other fluxing agent is used to facilitate the welding. It is very rarely that an exception occurs, and an ingot or a charge is discarded by the rail-mill. The mixture of pig-iron used for the production of this steel is melted in cupolas of very interesting construction (not to be de

Jan 1, 1874

By Lawrence Addicks

Some three years ago the smelter connected with the Chrome, N. J., refinery of the United States Metals Refining Co. found itself embarrassed by constantly increasing piles of unsmelted blast-furnace flue dust. The charge was fine and 10 to 15 per cent. of it was blown over into the flue. Sulphur was at that time too scarce to make the sintering of the finer part of the charge attractive, and raw-ore smelting in a reverberatory with a partly oxidized charge was not to be thought of. A great many schemes were considered, including leaching, blowing into the converters, and briquetting, the problem finally narrowing down to blast roasting and nodulizing. It was found by experiment that blast roasting would not yield a satisfactory sinter without the addition of coke dust or sulphide fines to augment the fuel value of the flue dust, and on this account, together with the low first cost of a kiln, the latter was decided upon. A couple of tons of flue dust were tried out in the experimental kiln in the laboratory at Yorktown, Va., of J. H. Payne, who acted as consulting engineer in this connection. This kiln was approximately 2 ft. in diameter and 20 ft. long and was fired with fuel oil. A run of several hours yielded excellent nodules. The ki1n showed no tendency to build in and form "nose rings" and gave a fuel consumption of 50 gal, of oil per ton of flue dust. There was no indication of any stack loss. The roasting was strongly oxidizing. A study of what full-size kilns were doing on cement clinker and on nodulizing pyrites cinder indicated that a 60 by 6 ft. kiln would certainly not consume over half of the oil per ton of flue dust shown on the test and as fuel oil at that time was selling around 2 1/2 c. per gallon it was decided to let it go at that. A 60 by 6 ft. kiln was decided on, as that size is standard in cement practice, although fast being replaced by much larger ones with their greater fuel economy, and as it fitted in the building space available. A 6-in. brick lining brought the net diameter down to 5 ft. The inclination

Jan 1, 1915

By Charles A. Kumke

THE "slot" system of mining in use at the Golden Queen mine, Mojave, Calif., does not involve any new mining methods. It is, however, a new combination and adaptation of several stoping systems in common practice. The first plans for stoping operations at the Golden Queen mine called for ordinary shrinkage stopes. The hanging wall of the vein, where exposed, is a very hard felsite with a dip varying from 50º to 70º. The vein itself is quartz, shattered and fractured enough to make it easy to drill and blast to a clean hanging wall. There is no well defined footwall. The footwall limits had to be determined by assays as the metal content faded out into the felsite. Two shrinkage stopes, each approximately 200 ft. long, were laid out and started on the 200-ft. level. One of these stopes, varying in width from 8 to 18 ft., was carried on up to the top limits of the ore, approximately 120 ft., without a break in the operation. The shrinkage system proved entirely satisfactory and low costs resulted from this operation. In the final drawing of this stope, two timbermen followed down on the ore, scaling the hanging wall clean of small fragments and placing an occasional long stull from foot to hanging wall where it was deemed necessary. Few such stulls, however, were required. The hanging wall stood up very well until the stope was entirely emptied of ore. Waste fill was then put into the stope by the glory-hole method, through a waste raise that had previously been run to surface from the middle of the stope. The second shrinkage stope ran into difficulties about 35 ft. above the sill floor. Up to this point the maximum width of the ore had been about 15 ft. The ore suddenly widened into the footwall to a distance of 40 ft. in the north end of the stope. In extending the stoping operations over this width, it was soon found that the vein material would not arch and support itself. Drilling in this end of the stope was then entirely too hazardous, and all operations were stopped. The ore was drawn through the chutes below as rapidly as possible and as soon as the stope was entirely

Jan 1, 1937

By W. J. Taylor

SOME eight years ago I became interested to a considerable extent in one of the well-known deposits of sulphury iron ore in New Jersey, and, as a consequence, soon became interested in the subject of the elimination of the sulphur. On investigation I found that no department of pig-iron manufacture had received less attention than that of roasting or calcining ores for the elimination of sulphur, carbonic acid, water, etc. My first devices to eliminate sulphur were naturally in the lime-kiln line, with solid fuel mixed with the ore, in the same way as common lime is burnt with small coal. While this process will succeed tolerably well for carbonates or calcareous ores, even though they contain sulphur as sulphates, I soon found it would not answer for our magnetites, which are very hard and dense, carrying from 2 to 5 per cent. of sulphur as pyrites, and become very sticky as soon as any attempt is made to roast them at a high heat. Some laboratory experiments developed the fact that the nearer the gaseous current surrounding the hot ore approached in composition pure atmospheric air the more rapidly the sulphur was oxidized, while in an atmosphere composed largely of carbonic oxide, carbonic acid, and nitrogen, it seemed almost impossible to reduce the sulphur more than one-half, which was a confirmation of the well-known fact that just one-half of the sulphur in pyrites volatilizes very easily, and that the trouble is mainly to remove the remainder. It was thus apparent that the atmosphere surrounding the ore was of as much importance as the temperature. This led to the trial of a kiln to heat the ore by conduction through the brick. It was divided into ore-chambers, which were surrounded with flues, through which the products of combustion of gaseous fuel passed without coming in contact with the ore. We did not, however, succeed by this arrangement in getting more than a cherry-red heat, which, of course, was too low to accomplish anything. I had at this point become convinced – 1st. That gaseous fuel only would accomplish the work successfully. 2d. That since a high temperature was required to insure a rapid oxidation of the sulphur, and also to make the ore open and porous,

Jan 1, 1881

By George R. Ensminger, E. Heaton Hemingway

During the past year, the Watertown Arsenal has been interested in the occluded gas and oxide content of certain ordnance steels in order to determine, if possible, whether some of the peculiar failures in their physical properties might not be traced to these causes. For the purpose of best accomplishing a study of this nature, it was decided to heat the various samples of steel in a vacuum for definite periods of time at high temperatures and carefully examine the quantity and composition of the gases evolved. By the use of a fused quartz tube fitted with a water-cooled, ground-glass stopper, it was possible to maintain a vacuum at a temperature of 1000" C. for 10 hr. without perceptible leakage being evidenced as measured by a McLeod gage. The silica tube was connected with an evacuating system so constructed that the volatilized gases could be removed from time to time and a low vacuum of 0.0002 to 0.0010 mm. of mercury continuously maintained. At no time during the period of heating or in the removal of the gases did the evacuated system come in contact with rubber stoppers or any rubber connections, making contamination from such a source, impossible. In other words, the only volatile products that could be: obtained must necessarily have come from the steel itself or from the glass and fused quartz system or from the mercury used as a sealing agent. In the carlier experiments, the steel treated was in the form of turnings, but this practice was discarded because the weight's of the turnings before and after being heated were widely inconsistent,, even when taken from the same piece of steel. This discrepancy developed largely through variation in the amount of metal volatilized with the gases; this metal could always be seen in the form of a ring where it condensed in the cooler portion of the tube in either the amorphous or crystalline condition. Because of the frequency in occurrence and the variation in the manner of deposition, this material was roughly analyzed and found to consist for the most part of iron. At the same time two samples of steel that had been

Jan 1, 1922

By E. Heaton Hemingway, George R. Ensminger

During the past year, the Watertown Arsenal has been interested in the occluded gas and oxide content of certain ordnance steels in order to determine, if possible, whether some of the peculiar failures in their physical properties might not be traced to these causes. For the purpose of best accomplishing a study of this nature, it was decided to heat the various samples of steel in a vacuum for definite periods of time at high temperatures and carefully examine the quantity and composition of the gases evolved. By the use of a fused quartz tube fitted with a water-cooled, ground-glass stopper, it was possible to maintain a vacuum at a temperature of 1000" C. for 10 hr. without perceptible leakage being evidenced as measured by a McLeod gage. The silica tube was connected with an evacuating system so constructed that the volatilized gases could be removed from time to time and a low vacuum of 0.0002 to 0.0010 mm. of mercury continuously maintained. At no time during the period of heating or in the removal of the gases did the evacuated system come in contact with rubber stoppers or any rubber connections, making contamination from such a source, impossible. In other words, the only volatile products that could be: obtained must necessarily have come from the steel itself or from the glass and fused quartz system or from the mercury used as a sealing agent. In the carlier experiments, the steel treated was in the form of turnings, but this practice was discarded because the weight's of the turnings before and after being heated were widely inconsistent,, even when taken from the same piece of steel. This discrepancy developed largely through variation in the amount of metal volatilized with the gases; this metal could always be seen in the form of a ring where it condensed in the cooler portion of the tube in either the amorphous or crystalline condition. Because of the frequency in occurrence and the variation in the manner of deposition, this material was roughly analyzed and found to consist for the most part of iron. At the same time two samples of steel that had been

Jan 1, 1922

By Peter R. Drummond

THE final casting of refined copper has been re-J- stricted for generations by the following sequence of operations: Filling the reverberatory furnace, melting, skimming, blowing or flapping, and poling. The hoped-for 24 hr cycle, producing 300 tons or more, has been taken up largely with the necessary bat time-consuming tasks of cleaning the bath, sulphur elimination, and in turn removal of excess oxygen to produce tough-pitch copper. Incidental to comparatively slow melting under combustion gases, copper oxides react with the furnace lining, and the slag so-formed must be completely recycled. The three-phase arc furnace has eliminated some of the cycle stages, and telescoped the remainder into a continuous operation. Electrical energy, supplied to graphite electrodes enclosed in high grade refractories, rapidly melts copper cathodes and sustains a stream of metal, containing approximately 0.01 pct oxygen, without contamination from fuel. The arc was struck on the first large electric furnace for melting copper in the United States on April 13, 1949. The earliest use of this type of furnace was at Copper Cliff, Ont., in 1936, and an admirable description of their installation has been published? Copper, melted in the Baltimore furnace, is used to cast billets, and the installation differs somewhat from the Canadian, as will be described. The arc furnace is a heavy-duty, three-phase furnace, holding 50 tons, the general outline of which appears on Fig. 1. The steel shell is 15 ft ID with a bottom radius of 14 ft 2 in. The roof, separate and distinct from the body, consists of a 15-ft water-cooled, cast-steel ring of the same outside diameter as the furnace. The center line of the furnace lies 9 ft 6 in. from that of the trunnions, permitting a 5" backward tilt for skimming, and a 40" maximum nose tilt forward for complete draining. Normally, the furnace overflows by displacement, and the use of the forward tilt arrangement is restricted to covering charging delays. The charging slot, 3 ft 8 in. x 5 in., lies on the north center line, the tap hole on the south, and the 30x30 in. skim door 45" to the west of the slot. The original 20-in. graphite electrodes were replaced with 14 in. in December 1949. Three conventional direct current winch drives, governed by electrical controls, position each electrode which has individual mast supports and counterweights. An independent circulation supplies cooling water for the electrode glands, the roof ring, charge slot, and the skim door frame. Arc Furnace Refractories Hearth: Fused-in monolithic bottoms had been used in copper arc furnaces, installed prior to April 1949. These consisted of thin layers of periclase, successively fused in place over preliminary brick courses. Heat was obtained from the arc, using a T-like arrangement of broken electrodes resting directly on the periclase to be fused. The operation, taking weeks to perform, was very expensive. The chemically-bonded magnesite-brick bottom, installed at Baltimore, was the first of its kind and a radical departure from previous practice. It consists of a 1 to 6-in. layer of castable refractory laid on the steel shell, modifying it to a 12 ft 2 in. bottom radius. Two courses of 9x2 % -in. fireclay straights and keys follow. The third course is made of 9-in. magnesite blocks of special shape to form circles of an inverted arch. It was started by a four piece keystone with skew-backs forming the outer course. The fourth course also started on a central keystone, or button, of four 90" segments, 12 in. diam x 13 Vz in. deep, and continued with 13%-in. blocks. Skewbacks at the shell completed the course to produce a horizontal surface for the side walls with a single course of No. 2 arch fireclay against the steel. Dry chrome-magnesite cement was brushed over each course after laying, and a 1-in. expansion space between the brick and the shell was filled with the same mixture. The total bottom thickness, excluding the castable material, was 5 in. of clay plus 22% in. of chemically-bonded magnesite. Tap Hole: A 5-in. OD and 3-in. ID silicon-carbide tube constitutes the tap hole and is set tangential to the upper course of the furnace bottom. Molten metal fills the tube when the furnace is level and filled to capacity. Side Walls: The lining, against the shell, consists of a 9x4Y2x3 in. soldier course of fireclay, using straights and No. 1 arches to turn the circles. A second soldier course of 9x4'/2x2'/2-in. fireclay was laid in a somewhat similar fashion. Three courses of 13Y2x6x3 in. and 9x6~3 in. of final magnesite, laid flat, completed the lining, using Nos. 1 and 2 keys to turn the circles. Cardboard spacers were placed between every two bricks in horizontal courses, and a thin coat of chrome-magnesite cement filled the joints between the firebrick and magnesite. A sprung-arch spanned the skim door with jambs of suitable magnesite shapes. Charge Slot: The slot is 3 ft 8 in. wide x 5 in. high. A silicon-carbide sill of special shapes has a 30" slope to allow cathodes to slide easily into the bath. The original arch was flat, and composed of Nos. 1 and 2 wedge magnesite with a 6-ft radius. It projected 5 in. over the sill, and, being a flat arch, gave an 18 15/16-in. opening between the inner edge and the metal line. The whole assembly was later raised 9 in., and the flat arch replaced with an arch, the lower edge of which maintained the 5-in, width from the outer to inner edges as shown in Fig. 2. A water-cooled, cast-copper jacket protects the steel shell behind the slot.

Jan 1, 1952

By M. Kaufman, A. E. Palty

The phase structure and aging characteristics of two nickel-base alloys, Inconel 718 and 702, were investigated. Wrought and cast Inconel 718 showed ArisCb as the major hardening phase, as well as ture was positively identified. Inconel 702 had Ni3Al as the major hardening phase, as well as Cr7C3-of the phase amounts with temperature was determined. 1 HE value of phase structure and aging rate information in the utilization of alloys has been well established. As part of a continuing program, the results obtained on two nickel-base alloys, hconel 718 and Inconel 702, are presented here. Inconel 718 is a relatively new alloy considered for application in both sheet and cast form. Its novelty lies in the use of high columbium tantalum levels in a Ni-Cr-Fe base. No previous work on this type of alloy has been reported in the literature. Inconel 702 is being used now as an oxidation-resistant sheet alloy. Occasional fabrication problems spurred work on its basic structural behavior. The composition of 702 is essentially nichrome plus aluminum and titanium for strengthening and added oxidation resistance. This puts it in the basic class of other nickel-base alloys which have already been studied. The two alloys will be treated separately. INCONEL 718 Phase Study Methods. The phases present after the various conditions investigated were determined by X-ray diffraction analysis of two different elec-trolytically extracted residues and microscopic examination. The extraction methods, using 10 pct HC1 in alcohol and 10 pct H3PO, in water and the X-ray technique have been described previously.' The effect of temperature on the phase behavior was studied on sheet material using the following heat treatments: 2250°F, 2 hr, ice brine quench 1300 100 hr, water quench 2250°F, 2 hr, ice brine quench 1400°, 100 hr, water quench 2250°F, 2 hr, ice brine quench 1500°F, 100 hr, water quench 2250o 2 hr, ice brine quench 1700°F, 48 hr, water quench 2250°F, 2 hr, ice brine quench 1850°, 24 hr, water quench 2250°, 2 hr, ice brine quench 2000°F, 6 hr, water quench The high solution temperature was an attempt to dissolve as many phases as possible without causing melting. Aging exposures were made longer than normal heat treating times in order to approach equilibrium conditions more closely. Cast material was examined only in the as-cast condition and with the normal heat treatment (1700°F, 1 hr, air cool 1325oF, 16 hr,air cool). Materials. The full phase study was performed on sheet specimens approximately 1/2 in. by 2 in. by 1/16 in. The cast pieces were cut from the gating system of a vacuum-cast test piece. An all weld-metal specimen was made by running many weld beads on the sheet using filler made from the sheared edges of the sheet. All external contamination was removed by belt-sanding followed by an electro-polishing treatment. The typical composition of wrought Inco 718 is as follows: The alloy is air-melted. Cast hco 718 may have higher molybdenum. Due to the expectation of finding a Ni-Cb phase (Ni, Cb) for which no X-ray diffraction pattern was available in the ASTM Index, a special 5 Ib heat to the stoichiometric composition of this intermetallic

Jan 1, 1962

By C. E. Brown, D. C. Helms, O. W. Shimer

The early history and progress of anthracite stripping, from the first known operation at Summit Hill in 1821 through 1917, was covered in 1917 in a paper by J. B. Warriner,1 then chief engineer, now president of the Lehigh Navigation Coal Co., and stripping methods between 1917 and 1931 were the subject of a paper in 1931 by H. H. Otto,2 of the Hudson Coal Company. . The present paper is confined to methods used today on pitching beds in the western middle and southern anthracite fields. The anthracite industry, not unlike other large industries, had been affected for more than a decade by the adverse economic conditions of the country. High development and operating costs and the competition of other fuels caused some collieries to be closed temporarily, while others were permanently abandoned. In order to obtain a larger tonnage at a lower development cost, a larger stripping program was put into effect by many companies. In 1932, stripping operations accounted for 3½ million tons, or 7 per cent of the total anthracite production of the year, and in the next 10 years this increased until more than double that amount, or over 14 per cent of the total production, was from stripping operations. These figures represent the total production of the entire anthracite region; unfortunately, the figures have not been separated from the different fields. There are some operations in the southern and middle western fields where the tonnage obtained from stripping amounts to more than 30 per cent of the total colliery production. Stripping operations on pitching beds present difficult problems. Owing to the various geological formations, a different situation is encountered in nearly every stripping pit; and because of the proximity of an overlying bed, it is very often necessary to load the entire overburden into trucks and haul it out of the pit to the dump area, although back filling and casting is done whenever possible. Beds in the areas described in this paper pitch from 20" to 90 and some are on the invert in many sharp folds and faults. Planning The more successful strippings have had adequate engineering analysis. Some work is being done without much preliminary planning, but many strippings would have failed completely if close supervision by the engineering department had not been . maintained. Locations of stripping areas are determined as follows: I. By constructing cross sections over the areas and drilling sufficient diamond-drill or well drill holes to study the strata and to determine the location, thickness

Jan 1, 1944

By D. W. Philips

The problem of strata control has of necessity been prominent in mining, and accounts of individual experience and research abound in the literature of most mining countries. Organized rescarch began in Britain by the formation of a committee, upon the recommendation of the Safety in Mines Research Board in 1923, to investigate methods of reducing accidents from falls of roof in coal mines. This committee examined current practice in supporting underground workings in the principal coal fields, and its findings have been published in the Board's publications.' - Toward the end of these investigations it was suggested by the Chief Inspector of Mines that committees of the several mining institutes be formed to examine the problem in their respective districts. The formation of these committees marked a new phase in research on strata control. The work, though being done in the separate coal fields, followed a scheme outlined by the chief Mining Engineer of the Research Board, and was coordinated by his office. FACTORS INVOLVED The subject is being considered in Britain from two aspects: 1. To provide. by supports, means to prevent falls of fractured ground. 2. TO develop methods whereby the strata can be maintained in the least broken condition and thus make falls less likely. The first aspect is the one that will most likely lead to immediate results from the accident standpoint. It depends largely upon the assimilation of already acquired knowledge regarding the properties of supports and their correct application. The second aspect is perhaps more fundamental, and involves also the properties of the strata and their behavior under the influence of the forces induced by mining. The problem involves many natural factors that cannot be controlled, such as depth, inclination and thickness of seam and nature of the rocks. Other factors, which may be considered to a certain extent controllable, cannot be varied readily and with sufficient frequency to facilitate easy investigation; some of these are: rate of advance of the working face, length of face, direction of the face with regard to cleat and inclination, system and arrangements of rock packing behind longwall faces. In any given case the very many factors present have a mutual influence upon the effects produced in the mine. For example, a study of the effects of rate of advance cannot be done without considering the effects of other factors such as nature of strata, direction of face and quality and arrangement of supports. Conversely, the length and direction of a working face

Jan 1, 1947

By D. W. Philips

The problem of strata control has of necessity been prominent in mining, and accounts of individual experience and research abound in the literature of most mining countries. Organized rescarch began in Britain by the formation of a committee, upon the recommendation of the Safety in Mines Research Board in 1923, to investigate methods of reducing accidents from falls of roof in coal mines. This committee examined current practice in supporting underground workings in the principal coal fields, and its findings have been published in the Board's publications.' - Toward the end of these investigations it was suggested by the Chief Inspector of Mines that committees of the several mining institutes be formed to examine the problem in their respective districts. The formation of these committees marked a new phase in research on strata control. The work, though being done in the separate coal fields, followed a scheme outlined by the chief Mining Engineer of the Research Board, and was coordinated by his office. FACTORS INVOLVED The subject is being considered in Britain from two aspects: 1. To provide. by supports, means to prevent falls of fractured ground. 2. TO develop methods whereby the strata can be maintained in the least broken condition and thus make falls less likely. The first aspect is the one that will most likely lead to immediate results from the accident standpoint. It depends largely upon the assimilation of already acquired knowledge regarding the properties of supports and their correct application. The second aspect is perhaps more fundamental, and involves also the properties of the strata and their behavior under the influence of the forces induced by mining. The problem involves many natural factors that cannot be controlled, such as depth, inclination and thickness of seam and nature of the rocks. Other factors, which may be considered to a certain extent controllable, cannot be varied readily and with sufficient frequency to facilitate easy investigation; some of these are: rate of advance of the working face, length of face, direction of the face with regard to cleat and inclination, system and arrangements of rock packing behind longwall faces. In any given case the very many factors present have a mutual influence upon the effects produced in the mine. For example, a study of the effects of rate of advance cannot be done without considering the effects of other factors such as nature of strata, direction of face and quality and arrangement of supports. Conversely, the length and direction of a working face

Jan 1, 1947

By J. Weertman

The paper of Gifkins and Snowden considers the interesting but difficult problem of determining the stress dependence of secondary (steady-state) creep at low stresses. These authors have concluded that at stresses below 250 psi (2 x 107 dynes per sq cm) the secondary creep rate of lead is proportional to the stress (viscous creep) and is not proportional to the stress raised to about a fifth power. The experimental data considered by them were obtained on tests conducted at room temperature and at 50°C. The lowest stress employed was 50 psi (3.5 X 106 dynes per sq cm). The authors pointed out the main difficulty in determining the stress dependence of creep at low stresses. The creep tests must be run for very long lengths of time. However they made no estimates of when an experimental creep rate determination must be rejected because it does not represent a true steady-state or minimum creep rate. In order to be certain that a creep curve is within the steady-state region, the total creep strain should be of the order of 0.1 to 0.2. For a creep test of a year's duration this requirement implies that a secondary creep rate smaller than about 10-3 per hr cannot be measured reliably. The corresponding creep rate for a 10-year test is 10-6 per hr. The creep rates of the tests that were considered by the authors to prove the existence of viscous creep were of the order of or less than 10-6 per hr. One can conclude reasonably that this data does not prove unambiguously that large strain steady-state creep rate of lead is proportional to the stress in the stress range of 50 to 250 psi (3.5 x 106 to 2 X 107 dynes per sq cm). Another technique can be used to obtain the stress dependence at low stress levels. The creep rate is a very sensitive function of temperature. The creep rate can be increased by very large amounts merely by increasing the temperature. We carried out steady-state creep tests on lead single crystals25 at temperatures up to 320°C. We were able to obtain creep rate data down to stresses as low as 35 psi (2.5 x 10' dynes per sq cm). Our smallest creep rate was 8 x 10-5 per hr. Thus we obtained large strain, steady-state creep rates to even lower stresses than were considered by Gifkins and Snowden. No evidence was seen for viscous creep. The creep rate was proportional to the stress raised to about a 4.5 power down to the lowest stresses. Since there is no reason to believe that changing the temperature should change the stress dependence of steady-state creep, we feel that large strain viscous creep does not occur in the stress range quoted by the authors for lead single crystals or large-grain polycrystalline samples of lead. This conclusion does not imply that viscous creep may nat occur in a lower stress range or in the same stress range for fine grain material or at creep strains very much smaller than 0.1. Support by the U.S. Office of Naval Research is acknowledged. Authors' Reply R. C. Gifkins and K. U. Snowden We thank Dr. Weertman for his discussion and although, as we hope to show, we do not agree with his reservations, we do concur in stressing the importance of ensuring that creep rates are reliably obtained. Dr. Weertman appears to be content to accept n = 1 for low stresses with fine-grained material but not for single crystals. We believe our results show that the former result cannot be accepted without also accepting the latter. We will also show that the probable errors in our minimum creep rates are insufficient to alter our conclusions, that the criterion proposed by Dr. Weertman is arbitrarily restrictive and his alternative experimental approach possibly invalid. 1) A principal result of our Fig. 1(a) is that n = 1 for polycrystalline specimens at room temperature and 50°C for stresses below -250 psi. There was evidence that crystal slip and grain boundary sliding contributed approximately equaily to the overall strain in this low-stress regime. This implies that either a) grain boundary sliding controls slip within the grains or b) both grain boundary sliding and crystal slip independently occur according to mechanisms which give n = 1. Alternative a does not seem acceptable, so we were forced to consider b. This led us to reexamine work on bicrystals by Strutt and Gifkins and plot curves Fig. l(b). Previously Strutt et al. (loc. cit.), had merged these points with others using the Zener-Holloman parameter and thus, we now believe, had been led to overlook the behavior where n = 1. Curves C and D in Fig. l(b) did appear to confirm the hypotheses that n = 1 for crystal slip at these low stresses and the sliding curve F was similarly of the expected form. It was comparatively easy to find a quantitative theory to account for n = 1 for sliding and the similarity of curves C and D to curve F (all obtained from the same set of specimens) led us to feel that the single-crystal curves were valid. 2) We believe the secondary creep rates for both the polycrystalline and single-crystal specimens to be in error by factors c2. In Fig. 6 creep curves for polycrystalline specimens of lead(1) and lead(II) are reproduced as curves a and b, respectively, and curve c is for a single crystal at 100 psi. It is clear that, although the attainment of secondary creep rate takes 2 years for a and 150 days for b, thereafter the curve is linear for periods of 7 and -1 year, respectively. The single crystal has a linear portion commencing after 20 and extending to 90 days. Creep extension was measured directly using a traveling microscope reading to 0.01 mm on gage lengths marked on the specimens; the gage lengths

Jan 1, 1968

By J. F. Bryson

Instantaneous outbursts of coal in underground workings have occurred frequently in various coal fields in Nova Scotia; British Columbia; Canada; South Staffordshire, England; and the states of Washington, Virginia, West Virginia and eastern Kentucky in the United States. Such outbursts in eastern Kentucky have been called "mountain bumps " because they are more likely to occur under mountain cover than elsewhere, but recently they have become generally known as "coal bumps." The eastern Kentucky coal field has been troubled by these outbursts since 1923. Many men have been injured and killed, and the morale of the workmen in the vicinity of such outbursts has been affected to such an extent that mine production has suffered for a considerable period following each outburst. Such phenonlenal occurrence in any coal field naturally has led to discussion as to the cause and means of prevention. In the eastern Kentucky field, local mine officials, workmen, mining engineers, and many others have offered many ideas as to the cause and many methods of prevention have been suggested, and tried without success. Some of the plans were either practically or economically unsound and in general impossible to follow. To enumerate all the suggestions would require a large volume. One cause mentioned was that they might be due to gas under high pressure, as in Lower Silesian mines, and the remedy reeommcnded was to drain the gas by drilling holes into a suspected area. A hole 30 ft. long was drilled into a 60-ft. pillar to allow the gas to drain off, hut the pillar bumped the next day. Realizing the importance of obtaining the best information available on coal bumps, the writer, with the permission of the Harlan County Coal Operators' Assoeiation, invited Mr. George S. Rire, Chief Mining Engineer of the U.S. Bureau of Mines, who had investigated bumps in several countries, to conduct an investigation of coal bumps in this field. Mr. Rice, by permission of the Director of the G.S. Bureau of Mines, accepted the invitation, and it was the writer's privilege, in company with

Jan 1, 1936

By J. J. Ormsbee

Attempts at coal-washing have been made in the southern states during the last twenty years; but it is only within the last four or five years that the practice has become at all general. It might perhaps be claimed as one of the blessings derived from our departed " booms;" for, during their sway, the supply of coal of all qualities, good and bad, could not equal the demand; but, with the subsidence of the inflated demand, came imperative calls for fuels of better quality; and washers, previously regarded as luxuries, became necessities. Among those now in use in this section are representatives of the following types or classes: the trough washer; the jig washer; the percussive table; and those washers in which a constant upward current of water effects the separation. Without having full statistics, it is safe to say that there are in successful operation in the South more washers of the last class than of any of the others. The purpose of these notes is to present data with regard to the constructiOn, operation, and results of one of these current-washers, based mainly on the plant at No. 2 Slope, Pratt Mines, Alabama. The coal is mined from the well-known Pratt seam, having here an average thickness of 3 feet 6 inches. It has distinct cleavage-planes, and breaks in cuboidal lumps; is bright black in color, firm in structure, and air-slacks only after considerable exposure. It burns freely, leaving a gray or buff-colored ash. The lump- and nut-coals are used for domestic and steampurposes (chiefly, however, for locomotive firing), and the slack for making coke. The specific gravity is 1.272. Analyses of Pratt Coal. I. 11. 111. IV. Min. Resources Authority. Phillips. McCalley. of U. s.,1892. Lupton. Fixed carbon,. . 67.90 61.600 64.30 63.82 Volatile material,.. 29.80 31.480 32.08 31.85 Moisture,.,...... 1.508 1.07 1.02 Ash,.... 2.30 5.416 2.08 3.31 Sulphur,. .. 0.83 0.918 0.47 0.70

Jan 1, 1896

By J. F. Bryson

Instantaneous outbursts of coal in underground workings have occurred frequently in various coal fields in Nova Scotia; British Columbia; Canada; South Staffordshire, England; and the states of Washington, Virginia, West Virginia and eastern Kentucky in the United States. Such outbursts in eastern Kentucky have been called "mountain bumps " because they are more likely to occur under mountain cover than elsewhere, but recently they have become generally known as "coal bumps." The eastern Kentucky coal field has been troubled by these outbursts since 1923. Many men have been injured and killed, and the morale of the workmen in the vicinity of such outbursts has been affected to such an extent that mine production has suffered for a considerable period following each outburst. Such phenonlenal occurrence in any coal field naturally has led to discussion as to the cause and means of prevention. In the eastern Kentucky field, local mine officials, workmen, mining engineers, and many others have offered many ideas as to the cause and many methods of prevention have been suggested, and tried without success. Some of the plans were either practically or economically unsound and in general impossible to follow. To enumerate all the suggestions would require a large volume. One cause mentioned was that they might be due to gas under high pressure, as in Lower Silesian mines, and the remedy reeommcnded was to drain the gas by drilling holes into a suspected area. A hole 30 ft. long was drilled into a 60-ft. pillar to allow the gas to drain off, hut the pillar bumped the next day. Realizing the importance of obtaining the best information available on coal bumps, the writer, with the permission of the Harlan County Coal Operators' Assoeiation, invited Mr. George S. Rire, Chief Mining Engineer of the U.S. Bureau of Mines, who had investigated bumps in several countries, to conduct an investigation of coal bumps in this field. Mr. Rice, by permission of the Director of the G.S. Bureau of Mines, accepted the invitation, and it was the writer's privilege, in company with

Jan 1, 1936

Andes Copper has an old camp and a new camp, both isolated until recent road improvements. The old camp is at Potrerillos and the new one at El Salvador, 25 road ,miles away. The company was incorporated in 1916 and the first mine, concentrator, leaching plant, smelter and town were built in the succeeding twelve years. The first blister copper was poured on January 24, 1927. Economic conditions determined the long development period but the end result was the design and construction of a mining complex that was all of a piece, unlike the piecemeal additions to El Teniente and Chuqui. The mining episode at Potrerillos came to an end in 1960 when the ore at La Mina became exhausted and the new mine, El Salvador was brought in. Both of these ore deposits come within four hpundred feet of the surface and had they been developed within the last five years might reasonably have been brought in as open pit mines. However, that is conjecture, and as it stands the operation at La Mina was so ideal for the mining method selected, blockcaving, it became a classic for the textbooks. On the other hand, El Salvador has turned out to be a blockcaving mine to test an engineer's skill. In mining the copper deposit at Indio Muerto mountain, there are ground conditions of every conceivable type. Tunneling in one section, the back can cave like sized pea gravel and only spiling can hold it; in another heading only the lifters are shot but often the rest of the round will sit down in one piece; elsewhere the bottom squeezes up and closes the drifts; and contrary to all other locations, there are sections where the block can be undercut without support. Today the scope of operations at Andes extends from mining to electrolytic copper. Ore is mined at El Salvador and the concentrator is also situated there. Concentrates pipelined 14 miles to Llanth, the water decanted, and then it is hauled by rail 40 miles and 5500 ft higher to the Potrerillos smelter and refinery. Potrerillos and El Salvador are respectively 90 and 70 miles from the port of

Jan 11, 1969

By M. H. Sommer

Many attempts have been made to develop a relation between the tensile properties and the deep-stamping qualities of metals commonly used in deep drawing. These operations are generally performed cold and the metal is always stressed beyond the elastic limit. However, the problem cannot be readily solved without a proper investigation of: (1) the relation between actual stress and actual strain in the plastic range and (2) the stresses that are set up in the metal during the process of deep stamping. Much work has been done in recent years to add to our knowledge on plastic deformation of metals and the shortcomings of the ordinary tensile test, particularly in the plastic range, have been sufficiently recognized. It has been widely acknowledged that the local contraction, which appears in a tensile test after the maximum load is reached, is not a general characteristic of the material but is peculiar to this form of testing. The plastic deformation of materials is explained by the slip of their crystals along certain planes and this rearrangement should be similar whether observed in a tensile, compression or torsion test, and if the results of any of these forms of testing show deviations, it must be due to other influences, as for example the local contraction in a tensile test. For this reason the tensile-test curve as obtained from the measurements taken on a tensile-test machine or registered by an autographic device has been supplemented by the so-called "actual stress-strain curve," which is constructed from the observed tensile-test data by eliminating the effect of local contraction and is in complete alignment with the stress-strain values obtained in a compresison or torsion test. The actual stress-strain curve shows the increase of actual unit stress in relation to unit elongation beyond the yield point and beyond the point of maximum load and indicates clearly the extent of strain-hardening which the ordinary tensile-test curve does not reveal. The following metals have been investigated in this paper and their actual stress-strain curves and, further, their char-

Jan 1, 1934

By C. J. Frosch, J. A. May, H. G. White, C. D. Thurmond

Gallium phosphide epitaxial layers were grown from the vapor phase on undoped single-crystal galliurn arsenide substrates in silica tubes by an open-tube wet-hydrogen process. The epitaxial layers were grown over a range of water pressures at three substrate temperatures. Excess donor concentrations were determined by surface barvier capacitance measurelrzents without removing the layers from the substrates. The excess dmlor concentration, ND-NA, is fo~ind to vary approxilnately inversely with the pressure of water added to the hydrogen carrier gas. This is the relationship that would be expected for singly ionized silicon donors on gallium sites in extrinsic galliunz phosphide, with the silicon coming from the SiO generated by the reaction of hydrogen with the silica tube. An increase in the partial pressure of water in the hydrogen stream decreases the SiO pressure. The results indicate that ni, the intrinsic hole and electrmt concentration for gallium phosphide at the three substrate temperatures, is smaller than the concentration estimated from available data for the density of states effective masses and the energy gap. Mass-spectrographic measurements confirm that the dono?, introduced into gallium phosphide is silicon. The equilibrium concentrations of silicon in vapor-flown gallium Phosphide have been estimated from available thernzodynamic information that includes the solubility measurements of silicon in gallium phosphide in equilibrium with a gallium-rich liquid phase. Satisfactory agreement with the measured silicon concentrations is obtained. FROSCH1 has described an open-tube process for growing single-crystal Gap from the vapor phase by a GazO transport mechanism. The method depends upon the reaction of H20 in an H2 carrier gas with a heated source of polycrystalline Gap which provides the necessary vapor species. When the temperature of these vapor species is lowered, super saturation occurs and single-crystal Gap will deposit on a suitable substrate. Unintentionally doped single crystals of Gap grown by the wet H2 process in silica tubes are n type. Evidence is presented to show that the donor introduced is silicon, and that a qua si-equilibrium model accounts for the inverse dependence of the donor concentration on the water partial pressure and predicts the magnitude of the donor concentrations. Ainslie et al. experimentally showed a similar inverse relationship between the carrier density and oxygen pressure for GaAs. Emission-spectrographic analyses showed a decrease in the silicon concentration with increasing oxygen overpressure for GaAs. Cochran and Foster suggested the theoretical possibility of suppressing silicon contamination by using Ga20 generated by the reaction of gallium with water vapor. 1) EXPERIMENTAL The apparatus and procedures are essentially the same as those described by Frosch.' The apparatus consists of a 25-mm-ID SiO2 tube extending through a controlled high-temperature flat zone for the location of the polycrystalline Gap source and a downstream temperature gradient falling at a rate of about 14°C per cm. The latter provides the region of super saturation for the location of the single-crystal substrate. The partial pressure of water in the inflowing hydrogen stream, pA2, O was controlled by mixing me-tered proportions of dry H2 with H2 saturated with H2O vapor at 0°C. The total gas flows were about 200 cu cm per min in all experiments. The Gap sources were prepared by pulverizing boat-grown polycrystalline ingots to pass a 20-mesh sieve. The substrates were cut from an undoped single-crystal boat-grown GaAs ingot purchased from Monsanto. This ingot had a carrier concentration of about 1015 atoms per cu cm, a resistivity of about 5 ohm-cm, and a mobility of about 5000 sq cm per v sec at 25°C. Substrates with dimensions of 1 by 1 by 5 x lo-' cm were employed. The growth faces were chemically polished (111) arsenic faces. Epitaxial layers, at least 7.5 x 10-3 cm thick, were grown,. This required from 1 to 24 hr depending upon the Pii2Q values and the temperatures. In all of the runs, the source temperatures were 50°C higher than the substrate temperatures. Samples were prepared

Jan 1, 1968