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By W. H. Howard

As the result of a series of analyses and volume determinations of gases discharged from the converters at the Garfield Smelting Co.'s. smelter at Garfield, Utah, it was found that a considerable amount of lead fume was escaping from the stacks of the converter plant, and `that if this lead could be recovered at a reasonable expense an attempt should be made to collect it. Two methods suggested themselves: (1) filtration through bags; (2) electrical precipitation of the condensed fume by a Cottrell plant. Filtration through Bags Tests were made in filtering converter gases through the ordinary woolen bags of lead-smelter bag houses to determine how long such material could withstand the corrosive action of the gases. While blowing on white metal only, the fabric was usually destroyed in 10 to 14 hr. In tests conducted in filtering gases when blowing on matte or slagging stage exclusively and when relatively considerable lead was being volatilized, bags in one instance lasted 60.5 hr. It therefore was evident that the Garfield converter gases did not contain in themselves sufficient neutralizing elements for the purpose of direct bag filtration, even during the slagging stage. Such results were more or less anticipated, but it was thus proved that under all conditions of blowing Garfield mattes a very considerable amount of neutralizing agent, zinc oxide or lime, or both combined, as in the Sprague process, would be required to render bag filtration possible. Again, owing to the constantly changing quantities of SO3 from a given battery of converters, depending on the number blowing at a given time and on the stages of the blow, it would be practically impossible to feed automatically the required amount of neutralizer. In order to safeguard the bags sufficient neutralizer would have to be continuously added to meet the maximum SO3 output, thus. wasting neutralizers and giving an undesirably diluted final product. Such a condition is entirely different

Jan 8, 1914

By G. D. Rieck, G. H. G. Vaessen, D. L. Vogel

Both doped and undoped tungsten single crystals exhibit hardness anisotropy with respect to the direction of testing in each of the crystallographic phes investigated. The hardness anisotropy of undoped single-crystal tungsten is rather pronounced in (001) and (110) planes but very minor in (111) planes. The anisotropy in (122) planes of the doped and drawn tungsten after recrystallization is in accordance with a previously published concept of a stringlike distribution of the doping material in the metal. ANISOTROPY in mechanical properties of single-crystal tungsten was investigated by Rose et al., Garlick and Probst,2 and Beardmore and Hull. 3 Although data on the Vickers microhardness of tungsten have been published (see, e.g., Boosz 4 and the references cited by him), no data on the microhardness anisotropy are available. We report, therefore, some of our own results on the microhardness and its anisotropy. The microhardness has been determined of the following samples. A) Polycrystalline tungsten samples taken from: 1) powder metallurgical tungsten "normally" doped with potassium-silicon-aluminum oxides (final contents inppm: K, 100; Si, 100; Al, 20; Cr, 20; Fe, 50; Mo, 400; Ni, 10) and in the form of: a) a 10-mm-diam swaged ingot annealed at 1600°C, the grain sizes varying between 15 and 250 p; b) a sheet rolled at about 800°C and recrystallized at 2000°C7 the grain sizes varying between 15 and 150 p; c) a 2-mm-diam swaged rod, as-worked, with elongated grains (average diameter of about 25 , lengths up to 1 mm); 2) undoped electron-beam-melted tungsten with very large grains (contaminations in ppm: K < 10; Fe < 20; Cr < 10; Cu < 0.1; Ni not detectable; inter-stitial~ not determined). B) Single-crystal samples taken from: 1) pure tungsten grown on a Pintsch wire by the van Arkel procedure (main contaminations inppm: Mo, 100; Si < 10; Cu and Fe, 2; A1 < 2; interstitials not determined); 2) zone-melted undoped tungsten (analysis as under A2); 3) large crystals in secondary recrystallized tungsten

Jan 1, 1969

By Dewitt T. Ring

Lack of reliable detailed production figures for earlier years, together with the loose nomenclature in reference to producing horizons and the application of the term "field" without any definite meaning or restriction, has made it impossible to gather statistical data suitable for the Institute&apos;s form. In common with all other producing areas in the United States, Ohio has felt severely the economic ills of the entire industry during the past year. With overproduction, not only of crude but also of refined products, going on throughout the country, the price structure as a direct consequence has weakened on Ohio oils. Valuable crude is being sold at a big loss today because producers arc willing to, and do, sell more than the market can absorb. The axe hangs heavy over the heads of Ohio producers, and will until something constructive is done to keep supply in proportion to demand. This not only applies to Ohio but to the entire industry. Production of Pennsylvania grade in Ohio was under 50 per cent proration, which has been in effect since Oct. 1, 1937. Three price cuts since February have lowered the posted price from $1.70 to $1.18, a 30 per cent reduction; other grades were less drastically reduced. Pipeline storage facilities were limited; many producers were faced with the problem of providing their own storage or the risk of ruining their wells. If they wanted to pull and abandon wells, there was no market for junk equipment. Many employees have been cut to half time and in a number of instances their services entirely dispensed with. Summary of Production and Drilling Both oil and gas production in the state during 1938 showed a definite decline over the previous year. (A table showing production by years from 1876 through 1937 was published in Vol. 127 of the Transactions.) The total production of crude oil in the state for 1938 was 3,357,914 bbl., as compared with 3,654,000 bbl. in 1937, a decrease of 7.1 per cent. At

Jan 1, 1939

By Dewitt T. Ring

Lack of reliable detailed production figures for earlier years, together with the loose nomenclature in reference to producing horizons and the application of the term "field" without any definite meaning or restriction, has made it impossible to gather statistical data suitable for the Institute&apos;s form. In common with all other producing areas in the United States, Ohio has felt severely the economic ills of the entire industry during the past year. With overproduction, not only of crude but also of refined products, going on throughout the country, the price structure as a direct consequence has weakened on Ohio oils. Valuable crude is being sold at a big loss today because producers arc willing to, and do, sell more than the market can absorb. The axe hangs heavy over the heads of Ohio producers, and will until something constructive is done to keep supply in proportion to demand. This not only applies to Ohio but to the entire industry. Production of Pennsylvania grade in Ohio was under 50 per cent proration, which has been in effect since Oct. 1, 1937. Three price cuts since February have lowered the posted price from $1.70 to $1.18, a 30 per cent reduction; other grades were less drastically reduced. Pipeline storage facilities were limited; many producers were faced with the problem of providing their own storage or the risk of ruining their wells. If they wanted to pull and abandon wells, there was no market for junk equipment. Many employees have been cut to half time and in a number of instances their services entirely dispensed with. Summary of Production and Drilling Both oil and gas production in the state during 1938 showed a definite decline over the previous year. (A table showing production by years from 1876 through 1937 was published in Vol. 127 of the Transactions.) The total production of crude oil in the state for 1938 was 3,357,914 bbl., as compared with 3,654,000 bbl. in 1937, a decrease of 7.1 per cent. At

Jan 1, 1939

By L. A. Morley, W. T. Parry

Engineering design problems encountered by mining engineers often depend on the properties of natural granular materials such as soil, poorly consolidated sediment, fault gouge, and hydrothermally produced clay. The physical properties of such granular materials frequently are determined by the properties of silicate clays. In most mining applications, the clays are in contact with water. Swelling of clay due to water infiltration can interfere with underground support and haulage systems. Water-saturated clay on fault, fracture, and bedding planes causes slope failures in open pit mines, roof falls and heaving of weak underclays in coal mines. Water may be removed and cementing agents may be produced in clays by electrical methods. When an electrical potential gradient is applied to a charged capillary by two electrodes, water moves toward the negative electrode.1 This electrokinetic phenomena is known as electroosmosis and has been used to successfully dewater clay minerals which lack strength because of water. A consolidation effect occurs when the flow of water is accompanied by a decrease in water content. Electrochemical changes in the clay result in cementing and irreversible hardening. Theory A mathematical basis of electroosmotic phenomena is presented by Helmholtz.3 Smoluchowski5 modifies Helmholtz’ model and presents Eq. 1 to describe fluid flow velocity due to electroosmosis. For this formula, the zeta potential can be measured experimentally. The Helmholtz-Smoluchowski model is reasonably valid for clay aggregates with large pores saturated with fresh water or dilute electrolyte solutions.° An assumption used for Eq. 1 (equations are given Table 1; terms are defined in Table 2) is that the thickness of the electrical double layer surrounding the clay particles is small compared with the pore radii. Because this is generally not the case, Schmid7 presents Eq. 2. The volume charge density in the pore may be approximated from cation exchange capacity and porosity measurements. Casagrande8 offers a semi-empirical relationship (Eq. 3) for volume rate of electroosmotic flow and finds that k, is relatively constant for most clays with an average

Jan 1, 1972

By James A. Barr

PUMICE and pumice are varieties of volcanic ejecta formed from lavas of the more silicic types. Chemical analyses of pumice, pumicite and perlite, from which synthetic pumice is now produced, will generally fall within the limits 65 to 75 pct silica, 9 to 20 pct alumina, less than 8 pct combined alkalies (sodium and potassium oxides), less than 3 pct calcium and magnesium oxide, less than 3 pct iron oxide, with smaller amounts of manganese, titanium, phosphorus and other oxides. Perlite is characterized chemically by having, in addition to the percentages of oxides given above, as high as 6 pct combined water. PROPERTIES Pumice is a natural silicic foam glass produced by volcanic eruptions. The molten glass foam, upon rapid cooling, trapped the steam and volcanic gases that caused it to remain extremely porous and vesicular in structure. Pumice is found in the form of large lumps several cubic feet in volume and also in grains only a few millimeters in diameter. Chemically, pumicite is of the same range of composition as pumice. The distinction between pumice and pumicite is one of particle size; the coarser materials are pumice; the fine particles, an appreciable percentage of which is in the range of 100 to 325-mesh or finer, are pumicite. If the erupting volcano ejects a silicic lava containing much gas and water vapor, the sudden release of pressure tends to free the gas. The viscosity of the lava, its rate of cooling and the amount of gas evolved, all have part in determining the degree of expansion of the glassy particles and the fineness of the vesicle. If the lava is of the correct viscosity and there is not too much gas, a pumice may be formed. Large amounts of gas may cause complete disruption into pumicite.30 In 1916, Barbour gave the name "pumicite" to the volcanic ash of Nebraska and since then the name has been widely used in industrial circles for the finer-grained commercial varieties. The glass of pumice and pumicite is transparent and colorless. It

Jan 1, 1949

By J. F. Crouch

INTRODUCTION Roll front uranium has unique geologic characteristics (L,) which require specialized mining procedures. Roll front uranium occurs in arkosic sandstones with the ore zone typically less than 100 feet in thickness with up to 400 feet of overburden. The areal extent of ore is often a few thousand acres. Within the ore zone, ore grade material occurs in C and S shaped rolls, which are a few feet in width and height but may run laterally for several hundred feet. Since the total mine plan may include mining all ore, a roll front uranium mine may cover a few thousand acres. For both economic and environmental reasons, a roll front deposit is mined as a series of phases or sub-pits (3, 4) as illustrated in Figure 1. A typical sub-pit is 400 to 500 feet across. Sub-pit 1 is stripped initially and the overburden is spoiled on the surface. Depending on haul distance and overburden characteristics, stripping is done with trucks and shovels (or front-end loaders) or scrapers. When ore is encountered in sub-pit 1, mining begins, and the stripping equipment is moved to sub-pit 2, with sub-pit 2 overburden also being spoiled on the surface. When ore is encountered in sub- pit 2, sub-pit 1 mining should ideally be complete so the mining equipment can move to sub-pit 2 and sub-pit 3 overburden can be spoiled in the mined out sub-pit 1 area. Ore is mined selectively in a series of 12 to 24 inch cuts by a 2-1/2 yard backhoe on the basis of radiometric probing. For effective drainage, ditches must be maintained on the boundary of sub-pits, and mining lifts are kept as level as possible, with the orebody slightly elevated as part of general good mining practice. Waste removal in the ore zone is done by self- loading scrapers near the orebodies and by push-pull or push loaded scrapers away from the orebodies. Each self-loading scraper cut is

Jan 1, 1983

By Daniel M. MacNeil, Carle R. Hayward, Raymond L. Presbrey

At the February, 1916, meeting of the Institute, a paper presented by Hayward and Raymond gave the results of a study on the effect of time in tempering medium-carbon steel, when the following conclusions were reached: It is not fair to consider the question settled when only three specimens were used for each treatment, but the results obtained indicate that for reheating quenched medium-carbon steel to temperatures below 500" it is only necessary to heat it through. Longer heating has little or no effect on the tensile strength, ductility, or hardness. At temperatures above 500°, increasing the time of treatment causes a slight falling off in hardness and tensile strength with a corresponding increase in ductility. In the discussion of the paper these conclusions were attacked and evidence was presented that seemed to throw doubt on their accuracy. The authors, however, showed that this evidence was based on high-carbon or alloy steels and no facts regarding medium-carbon steel were presented. During this study the heat treatment was carried out in a small electric resistance furnace. It held nine specimens, but when these were placed in the furnace the temperature was lowered to such an extent that 30 to 40 min. were required to reach the desired heat. The authors pointed to this as a weak point in the investigation and it seemed wise to supplement the research by work on similar steel but using a different method of heating. This has been done and the results are given in the present paper. The Rhode Island Tool Co. coöperated by furnishing the steel and turning the tensile specimens. The steel used in this investigation came in three bars 3/4 in. in diameter; analyses are as follows: Carbon. MaNGanese, Phosphorus, Sulfur Bar NUmber PeR Cent. Per Cent. Per Cent. Per Cent. 1 0.46 0.56 0.010 0.055 2 0.44 0.58 0.007 0.042 3 0.48 0.58 0.007 0.048

Jan 1, 1922

By Daniel M. MacNeil, Raymond L. Presbrey, Carle R. Hayward

At the February, 1916, meeting of the Institute, a paper presented by Hayward and Raymond gave the results of a study on the effect of time in tempering medium-carbon steel, when the following conclusions were reached: It is not fair to consider the question settled when only three specimens were used for each treatment, but the results obtained indicate that for reheating quenched medium-carbon steel to temperatures below 500" it is only necessary to heat it through. Longer heating has little or no effect on the tensile strength, ductility, or hardness. At temperatures above 500°, increasing the time of treatment causes a slight falling off in hardness and tensile strength with a corresponding increase in ductility. In the discussion of the paper these conclusions were attacked and evidence was presented that seemed to throw doubt on their accuracy. The authors, however, showed that this evidence was based on high-carbon or alloy steels and no facts regarding medium-carbon steel were presented. During this study the heat treatment was carried out in a small electric resistance furnace. It held nine specimens, but when these were placed in the furnace the temperature was lowered to such an extent that 30 to 40 min. were required to reach the desired heat. The authors pointed to this as a weak point in the investigation and it seemed wise to supplement the research by work on similar steel but using a different method of heating. This has been done and the results are given in the present paper. The Rhode Island Tool Co. coöperated by furnishing the steel and turning the tensile specimens. The steel used in this investigation came in three bars 3/4 in. in diameter; analyses are as follows: Carbon. MaNGanese, Phosphorus, Sulfur Bar NUmber PeR Cent. Per Cent. Per Cent. Per Cent. 1 0.46 0.56 0.010 0.055 2 0.44 0.58 0.007 0.042 3 0.48 0.58 0.007 0.048

Jan 1, 1922

By E. H. Crabtree, Neil S. Parker

THE custom milling plant of the Eagle-Picher Mining and Smelting Co. is at Sahuarita, Arizona, approximately 20 miles south of Tucson. It is connected by a 2-mile railroad spur to the main line of the Southern Pacific Railroad between Tucson and Nogales. The capacity of the plant is 500 tons per day. The mill is conventional except perhaps in the rather wide variety of ores treated. These vary from straight copper ores, from which only one concentrate is produced, and lead-zinc or copper-zinc ores from which two concentrates are produced, to complex gold-silver-copper-lead-zinc ores, which require the simultaneous production of three or four concentrates. Custom ores are received in lots varying in size from a few truckloads to several carloads. Similar ores, after crushing and sampling, may be commingled, or individual lots may be milled individually. The ores are first dumped either into one of two 200-ton railroad bins, or into one of the two 75-ton truck bins. From these bins the lot of ore is individually crushed in a 24 by 16-in. jaw crusher to 3-in. size, and then to 1/2-in. size in a 3-ft Symons shorthead crusher. After crushing to 1/2-in., the ore is conveyed to the automatic sampling plant. This consists of three Vezin samplers in series, each successively cutting out 10 pct, to pct and 5 pct of its feed. After No. 1 sampler, the material passes through a mixing barrel and is then crushed to 1/4-in. in a set of rolls before passing to No. 2 sampler. Between No. 2 and No. 3 samplers, the ore is again mixed in a mixing barrel. The final sample, consisting of one pound per ton of ore, is taken to the laboratory, where it is further crushed to 1/8-in. in a coffeemill, and then riffled down to the size required for grinding in a pulverizer. The capacity of the crushing and sampling plant is up to too tons per hour, depending upon the character of the ore. After sampling, the ore is stored in five ore-storage bins, consisting of four 200-ton bins and one 175-ton bin. Each of these bins is discharged by means of belt feeders and conveyor to the ball mills, so that ores may be composited for milling in any required ratio, or milled separately by themselves. Fine grinding is done in one 6 by 4-ft Allis-Chalmers ball mill and one 8-ft by 36-in. Hardinge mill. These two mills both discharge to one 54-in. Akins Simplex Highweir classifier. The classifier sands can be returned to either or both mills, so that flexibility in grinding for different ores is easily obtained. Flotation equipment consists of four 66-in. Fagergren machines for copper-lead roughing and six similar machines for zinc roughing. Lead-copper bulk concentrate is cleaned and recleaned in eight No. 18-S Denver machines, and zinc concentrate is cleaned in three 66-in. Fagergrens. Copper-lead differential separation

Jan 1, 1947

By F. C. Dyer, H. E. T. Haultain

There has been much written on ball-mills, but no small amount of the literature is simply the expression of individual opinion without sufficient data. This is no doubt due to the complexity and obscurity of the phenomena involved. The outstanding paper is by E. W. Davis,I superintendent, School of Mines Experiment Station, University of Minnesota, which is based on a large amount of data from careful experiments with full-size apparatus, together with some mathematical philosophy. The entire philosophy is summed up in the equations given; but the present writers question the foundation upon which this philosophy is based. Encouraged by the work of Davis, the authors of this paper continued work they had been doing along this line. At first, photographs were taken of the ends of rods in a small laboratory rod mill. The mill was tilted slightly toward the feed end and the discharge end was removed. The rods remained in the mill long enough to permit good observation and good photography. Later, a mill 24 in. inside the stepped lining was constructed. It consisted of two sheets of plate glass, 1/2 in. apart, set in a cast-iron ring. It was driven by a variable speed motor and was connected to an accurate speedometer (magneto and voltmeter type). In place of balls or rods, cast brass disks were used. These were 3/8 in. thick and varied in diameter from 2 in. down to 3/4 in. Four small projections on each side prevented the disks sticking to the glass when water was used; sometimes steel balls % in. in diameter were added to the load of disks. There was also constructed a 6-in. mill with corrugated linoleum lining and containing a dry load of various sizes of steel balls. The mills were so connected to the motor that the small mill made two revolutions to one of the large mill. Many runs were made with a great variety of materials, including crushed marble, seeds of many kinds, and many mixtures (some wet and

Jan 1, 1923

By F. C. Dyer, H. E. T. Haultain

There has been much written on ball-mills, but no small amount of the literature is simply the expression of individual opinion without sufficient data. This is no doubt due to the complexity and obscurity of the phenomena involved. The outstanding paper is by E. W. Davis,I superintendent, School of Mines Experiment Station, University of Minnesota, which is based on a large amount of data from careful experiments with full-size apparatus, together with some mathematical philosophy. The entire philosophy is summed up in the equations given; but the present writers question the foundation upon which this philosophy is based. Encouraged by the work of Davis, the authors of this paper continued work they had been doing along this line. At first, photographs were taken of the ends of rods in a small laboratory rod mill. The mill was tilted slightly toward the feed end and the discharge end was removed. The rods remained in the mill long enough to permit good observation and good photography. Later, a mill 24 in. inside the stepped lining was constructed. It consisted of two sheets of plate glass, 1/2 in. apart, set in a cast-iron ring. It was driven by a variable speed motor and was connected to an accurate speedometer (magneto and voltmeter type). In place of balls or rods, cast brass disks were used. These were 3/8 in. thick and varied in diameter from 2 in. down to 3/4 in. Four small projections on each side prevented the disks sticking to the glass when water was used; sometimes steel balls % in. in diameter were added to the load of disks. There was also constructed a 6-in. mill with corrugated linoleum lining and containing a dry load of various sizes of steel balls. The mills were so connected to the motor that the small mill made two revolutions to one of the large mill. Many runs were made with a great variety of materials, including crushed marble, seeds of many kinds, and many mixtures (some wet and

Jan 1, 1923

By Frederic Laist

IN the spring of 1931 an experiment was made at Anaconda with a small reverberatory furnace, with which was combined a roasting shaft. The arrangement is shown in the accompanying drawing (Fig. 1) and requires but little explanation. The general idea was to roast .flotation concentrates, mixed with the proper fluxing materials, by dropping them down a heated shaft, from which the calcines would enter the smelting zone of the reverberatory furnace through an opening in the roof. The hearth of the reverberatory was 3 ft. wide by 21 ft. long. The roasting shaft was 3 ft. square inside and had an effective height of about 20 ft. . The reverberatory furnace was arranged so that it could be charged either directly through the roof with regular hot calcines, in the usual manner, or by means of the shaft, at the top of which were a feed-ing arrangement by which cold flotation concentrates, suitably fluxed, could be fed continuously and a pipe through which cold air was blown in. Means were provided for distributing the feed over the cross-section of the roasting shaft. At first a mechanical arrangement was used but later it was found that a simple cone served the purpose equally well and gave less operating trouble. On account of the small size of the equipment, it was realized that absolute capacity and fuel consumption figures would mean but little. It was not to be expected that ratios comparable to those obtained in large furnaces could be gotten, and the only way in which any idea could be obtained as to the possibilities of the combination furnace was to run it in parallel with the reverberatory furnace charged in the usual manner and operating without the shaft. The results of such parallel runs are given in Table 1. The amount of oil burned in the reverberatory was purposely maintained at substantially the same rate during both runs in order that the benefits resulting from the shaft might be gaged.

Jan 1, 1933

By Edward Keller

GOLD MOVEMENT ON SURFACE OF AURIFEROUS COPPER WHEN LATTER IS SUBJECTED TO OXIDATION SOME years ago the writer&apos;s attention was called to the fact that rolling-mill scales from auriferous copper do not have the gold content proportional to the gold contained in the copper from which they are derived. I So, later, he made a few simple experiments, which are described herewith. In: order to test the gold movement on the changing surface of a copper plate under the influence of oxidation, plates 6 in. by 8 in. by I in. (15.2 cm. by 20.3 cm. by 2.5 cm.) of converter copper, refined, of 99.29 per cent. purity, and leaded copper with a content of 95.65 per cent. copper, were obtained. Ten cuts were made through the thickness of the plates along the longitudinal center and the silver and gold content ascertained in each of the 10 samples in both plates, as shown in the accompanying illustration. The assay results, recorded in Tables 1 and 2, show that the gold distribution is relatively uniform in both plates, while silver has greater variations, especially in the leaded plate. The plates subjected to oxidation were placed in a red-hot muffle for about 2 hr., two small scorifiers being used as supports. Arranged alongside each other, they nearly filled the width of the muffle, and so were practically under equal conditions of heat and draft. The accurate observance of the time was not possible, however, because of other work, but the time of the exposure of each plate in a pair was identical. At the end of each period, each plate was plunged into water and completely freed of scale, which was carefully collected, dried, weighed, and analyzed. Samples of the plates had also been taken of the molten metal and duly analyzed. Table 3 contains the analytical data for the two plates and their several sheets, or layers, of oxide scales. These figures show that in the oxidations of the converter copper the silver is fairly uniform, while

Jan 1, 1919

By F. W. Jr. Sperr

THE byproduct coke oven is the most important artificial source of fuels for metals melting. Its products are solid, liquid, and gaseous in form. The amount of coke and primary byproducts obtained per ton of coal varies with the nature of the coal. The following represents the results obtained in many plants: Metallurgical coke (over 3/4 in.) 69 per cent. of coal Domestic coke (1/2 to 3/4 in.) 2 per cent. of coal Breeze (under 1/2 in.) 4 per cent. of coal Surplus gas, cubic feet per net ton 6600 Gas for heating ovens, cubic feet per net ton 4400 Tar, gallons per net ton 9 Ammonium sulfate, pounds per net ton 25 Benzols, gallons per net ton 4 All of these products, except ammonium sulfate, have important fuel value. Part or all of the metallurgical coke may be crushed or screened to furnish additional domestic coke. If producer gas is used, there will be a surplus of 11,000 cu. ft. of coke-oven gas per net ton of coal. The use of coke in the metallurgical industries has been retarded by the mental attitude of writers who disregard the dependence of coke quality on the conditions of its manufacture. Terms like "byproduct coke,"&apos; "beehive coke," "48-hour coke," and "72-hour coke" are only generally descriptive and tell nothing of the coke quality. For example, one writer1 stated that in brass-foundry practice, Connellsville coke is considered much superior to byproduct coke on account of the high sulfur content of the latter. Another2 says, "At equal prices for hard Connellsville coke and for byproduct coke at the foundry door, the hard coke is unquestionably cheaper. It will melt more metal per ton than the byproduct coke and requires less labor, as the byproduct coke is less dense and burns faster, thus requiring more frequent firing." A little study of the conditions under which coke is manufactured will show the error in mak-

Jan 10, 1920

By Robert W. Hunt

I have frequently stated that while the chemical composition of steel is important, yet even greater importance is connected with the mechanical and heat treatment of the metal. During the past year I encountered such a positive example of that fact that I deem it worth putting upon record. A prominent railway system divided an order for open-hearth steel rails between two steel works, both of which are under the control of the same corporation; giving to one about 18,000 tons, and to the other 7,500 tons of the same section and to be made under the same specifications. The rolling results obtained in the two mills varied so widely that a study of the figures is intensely interesting and serves to illustrate the advantages obtained by carefu1 ingot casting, and, quite as pertinently, the possible benefits of careful subsequent heating and rolling with moderate reductions in the rolling process. The smaller order was rolled complete in four installments during the same months that the larger order was being made in six separate installments; but, to permit of exact comparisons, the totals of the four installments (completing the order with the mill which I will call A) are given with the totals of the first four installments of the larger order, made by the mill designated as B. These rolling results are: Mill A. Mill B. Number of rollings,............................................ 4 4 Total number of rails rolled................................ 18,278 27,832 Percentage of rails cut to short lengths because of flaws near ends, etc......................................... 0.7 7.6 Percentage of rails made second quality for flaws, etc. 0.9 6.3 Percentage of rails scrapped for flaws, etc............... 0.6 4.5 Percentage of rails scrapped for failure at drop test... none 3.4 It should be stated that the figures given above cover all the reasons for putting the rails in the classes stated, Thus a part of the percentage of rails cut to short lengths may have been

Jan 1, 1914

By L. I. Cothern

The introduction of mechanized mining has created a demand for long working faces. It has also prompted mining men to contest the old theory that longwall methods can be used only where roof conditions are ideal and where pack-wall material is plentiful. "Longwall" as used in this paper means "modified longwall," "semi-longwall," or any of the various &apos;&apos;wall" or "long-face" methods of mining. Many mines are now using long conveyor faces that differ only in their methods of development, types of roof support and method of loading coal onto the conveyor. Some types of longwall operations are being carried on successfully under conditions heretofore considered impossible by men experienced in longwall mining. With possibly one or two exceptions, long-face methods in Alabama are a direct result of the need for lower mining costs. The advantages over room-and-pillar systems were clearly expressed in an article by Messrs. Fies and Lacy of the DeBardeleben Coal Corporation1 as follows: (1) increased tonnage per miner; (2) increased percentage of lump coal; (3) reduction in cost of development; (4) elimination of the cost of room yardage; (5) reduction in the cost of haulage, from ability to use larger cars; (6) reduction in the cost of machine cutting; (7) reduction in cost for tracks; (8) reduction in the cost of deadwork; (9) reduction in the cost of supervision. An attempt on the part of the operator to gain these advantages by changing from room-and-pillar to longwall requires considerable courage, and credit should be given to those men who have done their bit to meet the competition of rival fuels. The initiatory problems to be solved are in reference to: (1) retreating or advancing system, (2) length of wall, (3) method of roof support, (4) method of entry development, (5) type of conveyor or loader, (6) hand vs. machine loading (onto conveyors), (7) hand vs. machine loading in entries, (8) control of roof pressure,

Jan 1, 1940

By L. I. Cothern

The introduction of mechanized mining has created a demand for long working faces. It has also prompted mining men to contest the old theory that longwall methods can be used only where roof conditions are ideal and where pack-wall material is plentiful. "Longwall" as used in this paper means "modified longwall," "semi-longwall," or any of the various &apos;&apos;wall" or "long-face" methods of mining. Many mines are now using long conveyor faces that differ only in their methods of development, types of roof support and method of loading coal onto the conveyor. Some types of longwall operations are being carried on successfully under conditions heretofore considered impossible by men experienced in longwall mining. With possibly one or two exceptions, long-face methods in Alabama are a direct result of the need for lower mining costs. The advantages over room-and-pillar systems were clearly expressed in an article by Messrs. Fies and Lacy of the DeBardeleben Coal Corporation1 as follows: (1) increased tonnage per miner; (2) increased percentage of lump coal; (3) reduction in cost of development; (4) elimination of the cost of room yardage; (5) reduction in the cost of haulage, from ability to use larger cars; (6) reduction in the cost of machine cutting; (7) reduction in cost for tracks; (8) reduction in the cost of deadwork; (9) reduction in the cost of supervision. An attempt on the part of the operator to gain these advantages by changing from room-and-pillar to longwall requires considerable courage, and credit should be given to those men who have done their bit to meet the competition of rival fuels. The initiatory problems to be solved are in reference to: (1) retreating or advancing system, (2) length of wall, (3) method of roof support, (4) method of entry development, (5) type of conveyor or loader, (6) hand vs. machine loading (onto conveyors), (7) hand vs. machine loading in entries, (8) control of roof pressure,

Jan 1, 1940

By D. T. MacRoberts

A simple disturbance in a gas reservoir travels with a finite velocity which is nearly independent of the amplitude of the disturbance. As a result very complex transients may be set up which seriously affect observation at a well. The time of transient is much longer than commonly supposed. INTRODUCTION A change in pressure at a gas well is not propagated through the reservoir instantaneously but with a definite fin- ite velocity. Knowledge of the period ! of this transient is of value because: 1. Observations of pressure and flow-rate during the transient may affect interpretation of back-pressure tests. 2. Interference effects between wells may be used to determine continuity of formation. 3. Additional information concerning the physical characteris- tics may possibly be deduced from the behavior of a well during the transient. MATHEMATICAL DEVELOPMENT It will be assumed that the well is initially closed in and that the pressure is constant throlughout the reservoir. As the initial disturbance passes outward from the well it is further assumed that a logarithmic pressure distribution is established between the most distant point reached and the well, so that: <l> = ((/)r-<t>w) I ------------ + (/) . (1) Manuscript received at office of the Branch September 15. 1948. Paper presented at Branch Fall Meeting, Dallas. Texas. Oct. 4-6, 1948. r. = distance to any point between r and rw r = radius of drainge iw= radius of well (/) = potential (P for incompressible fluids, P2 for perfect gases) Up to the time the effect reaches the point r the production from the well will be: W= (Pr-P,,,) . (2) i b W = volume produced t = sand thickness F = fractional porosity Pr = reservoir pressure Pm = average pressure between r and rw Pb = atmospheric Pressure All pressures will subsequently be stated in atmospheres and the constant Pb will be omitted. This will be produced at a rate: dw t k (Pr³-PW2) do d 0 - . (3) log — Pw = pressure at the well k = permeability 0 = time = viscosity

Jan 1, 1949

By D. T. MacRoberts

A simple disturbance in a gas reservoir travels with a finite velocity which is nearly independent of the amplitude of the disturbance. As a result very complex transients may be set up which seriously affect observation at a well. The time of transient is much longer than commonly supposed. INTRODUCTION A change in pressure at a gas well is not propagated through the reservoir instantaneously but with a definite fin- ite velocity. Knowledge of the period ! of this transient is of value because: 1. Observations of pressure and flow-rate during the transient may affect interpretation of back-pressure tests. 2. Interference effects between wells may be used to determine continuity of formation. 3. Additional information concerning the physical characteris- tics may possibly be deduced from the behavior of a well during the transient. MATHEMATICAL DEVELOPMENT It will be assumed that the well is initially closed in and that the pressure is constant throlughout the reservoir. As the initial disturbance passes outward from the well it is further assumed that a logarithmic pressure distribution is established between the most distant point reached and the well, so that: <l> = ((/)r-<t>w) I ------------ + (/) . (1) Manuscript received at office of the Branch September 15. 1948. Paper presented at Branch Fall Meeting, Dallas. Texas. Oct. 4-6, 1948. r. = distance to any point between r and rw r = radius of drainge iw= radius of well (/) = potential (P for incompressible fluids, P2 for perfect gases) Up to the time the effect reaches the point r the production from the well will be: W= (Pr-P,,,) . (2) i b W = volume produced t = sand thickness F = fractional porosity Pr = reservoir pressure Pm = average pressure between r and rw Pb = atmospheric Pressure All pressures will subsequently be stated in atmospheres and the constant Pb will be omitted. This will be produced at a rate: dw t k (Pr³-PW2) do d 0 - . (3) log — Pw = pressure at the well k = permeability 0 = time = viscosity

Jan 1, 1949