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By John M. Hartman

TWO systems are used for heating air in blast-furnace operations: I. The double surface system, in which a cast-iron pipe is heated on the outer surface, and, at the same time, heats the blast from its inner surface. This is simple in operation and gives a continuous effect, but is limited by 1100" F. as a maximum temperature of the blast. 11. The single surface system, by which large surfaces of fire-brick are heated, and air passed over the heated surface, absorbing the heat and carrying it on to the furnace. This system is more comylex than the double surface, as it invoives the reversing of the air and gas every hour and a half. The single surface system has two advantages: 1. The blast can be heated to a temperature of 1800° F. 2. The stoves are indestructible. From recent experience it has been found that 1300" F. to 1400' F. is the best average temperature for economy of coal for safe working. This is equivalent to a saving of 11/2 to 2 cwt. of coal per ton of iron over the extreme limit of cast-iron stoves. Independently of this, it is a strong point in favor of this system that the blast can be raised to a temperature of 1800° F. within an hour when the hearth is getting cold. All furnacemen know the ralue of a hot hearth for quality and quantity of iron. Cooling of a hearth occurs from leaky tuyeres, scaffolds, or heavy burden. When there is not sufficient coal at the tuyeres to seize on the oxygen of the entering air and convert it at once to carbonic oxide, there will not be enough heat to liquefy the cinder. Black cinder and poor iron are the results. The remedy is additional heat from the blast. If the stoves will give 1800" in place of 1100°, it is obvions that the furnaces will get around sooner, and without waiting for a change of burden at the tunnel head to bring extra coal to the hearth. The heat absorption caused by a leaky tuyere will chill the hearth and drive the zone of fusion higher up in the furnace. This loss must be supplied, and the calorics lost to the hearth must be regained before good iron can be made. Take another case: A furnace carefully burdened on No. 1 iron, during a spell of damp weather, goes on to No. 2 or No. 3. Heat is lost to the furnace through absorption by the moisture, and less burden must be carried in order to get back on No. 1. When the

Jan 1, 1879

By R. N. Orava, L. A. Rice

One of the difficulties encountered in the preparation of titanium specimens for transmission electron microscopy is the formation of a surface hydride phase during the thinning operation at ambient temperatures. This necessarily destroys contrast, obscures features of interest, and often leads to misleading results. By cooling the polishing solution to temperatures below 0°C this problem has been overcome in the past.' However, since the cooling procedure is somewhat inconvenient, a technique was developed at this laboratory for thinning titanium at room temperature. The material examined was A-70 grade commercial-purity titanium purchased from the Crucible Steel Co. of America. The mechanical and thermal treatments, described elsewhere,* yielded a recrystallized grain size of 4 to 5 p. A slice, 0.5 mm thick and 1.5 mm wide, was cut from a bulk specimen2 using a diamond cut-off wheel. The sample was then coated with paraffin wax. Two concentric holes, about 0.03 in. in diam, one at each broad face, were scribed through the wax to the surface of the metal. Subsequent polishing for several minutes resulted in a preferentially thinned or "dimpled" region. Thereafter, the wax was removed and the entire specimen polished to the point of breakthrough at the base of the dimple. Polishing was effected chemically in a solution comprised of 30 pct HF and 70 pct HN03 by volume, kept at room temperature (20" to 25OC). The specimens were mounted in a stereo holder and examined at 100 kv in a Philips l00 B Electron Microscope. A micrograph from a foil prepared in this way is shown in Fig. 1. Several points are noteworthy. This electron micrograph shows neither evidence for the formation of a hydride phase due to specimen preparation nor, moreover, any obvious surface contamination in electron-transparent areas of a freshly prepared foil. Localized variations in foil thickness observed in some specimens seem to be a characteristic of this chemical polish but do not seriously detract from foil quality. With care, the lifetime of a good specimen exceeded 2 hr in the beam. On the other hand, the exposure of a sample to air at atmospheric pressure, when still warm from the beam, clearly resulted in the deterioration of a formerly clean area just previously in the field of view. An example in Fig. 2 of a contaminated region reveals features, and "spots", not unlike those observed in much purer material by Yamane and Ueda. Consequently, one can conclude that: 1) titanium can be thinned satisfactorily in a solution at room temperature for transmission electron microscopy; 2) large-scale precipitates, >50A in diam, are absent in commercial-purity, recrystallized, furnace-cooled, A-70 titanium containing about 3800 wt ppm interstitials and the same amount of iron; 3) care must be taken not to attribute surface contaminants to precipitates. The authors are grateful to the Franklin Institute for financial support and permission to publish.

Jan 1, 1967

By James J. Wise

INTRODUCTION One basic design feature utilized in all caving and sublevel stoping techniques is the drawpoint. These draw- points may be located on sublevels when using sublevel caving methods, or may be located at the floor pillar when using block caving, sublevel stoping, or vertical crater retreat methods. How the drawpoint is designed and developed, how the ore is withdrawn from it, and how the ore is hauled to the shaft or portal encompasses the loading and hauling function. Historically, predominate loading and hauling equipment for caving and sublevel stoping has been scraper slushes, air powered overshot loaders, and rail mounted train haulage. All of this equipment has two common constraints. 1. The equipment locked into specific work sites and is not easily shifted to another work site. This constraint can often lead to low utilization and low productivity of the equipment. 2. The equipment is designed to work in the horizontal plane, which requires that the various horizontally-oriented working levels be connected by vertical raises. Vertical development by means of raises is not only increasingly expensive, but also is not well adapted to movement of large equipment into the s topes. In the last 20 years, two revolutionary labor saving concepts have become common in caving and sublevel stoping. 1. Long-hole drilling, which has mechanized the development of the drawpoint extraction structures as well as the actual stoping and undercutting. 2. Vertical access by means of trackless ramps between levels, and the consequent use of rubber-tired vehicles not only for loading and hauling ore, but also for the rapid movement of men, material, and equipment into the stopes. The introduction of trackless mining systems for loading and hauling has not been easy. Mine planners have had to re-orient their thinking to the larger operating space required by trackless equipment. Larger access drifts have necessitated reconsideration of mine layouts to insure proper ground support. Spiral ramp developments were often laid out on slopes steeper than the equipment could negotiate economically. Mine operators had to accept the added responsibility of roadway maintenance, water removal, increased ventilation, and more technical mechanical maintenance. As the mine operators enthusiastically embraced the superior productivity of trackless mining systems, they some- times disregarded opportunities to combine the older types of equipment with trackless vehicles to maximize productivity at lowest cost. In order to gain a proper perspective on this problem, it is necessary to review the various techniques of loading and hauling and to become more aware of the advantages or disadvantages that each system or combination contributes. LOADING Loading consists of two functions, (1) primary loading from a drawpoint and (2) secondary loading into long- haul system, a truck, train, or possibly a belt conveyor. Mine planners have four basic loading systems available to fit various operating conditions. These are, (1) gravity, (2) scrapers, (3) rail-mounted loaders, and (4) trackless loaders. Each system can produce optimum results under specific conditions. Gravity Loading Gravity loading is probably the oldest system in general use today.

Jan 1, 1981

By T. P. Papazoglou, N. A. D. Parlee, W. C. Phelps

PRACTICAL designs for good "home made" molybdenum furnaces are hard to find in the literature. The one described briefly below left something to be desired but was good enough to operate as a rather long (30 in.) two-zone furnace for periods of over 2 months at temperature (1400° to 1600°C) without shutdown. The principal problem in our application was to establish a temperature profile with two distinct constant-temperature zones and no dip between the zones. This and other problems mentioned below would not arise, however, if both zones were operated at the same temperature and if a shorter furnace were acceptable. Since molybdenum windings oxidize very rapidly when heated in air, impervious refractory tubes were used for both the furnace core and the outer jacket tube shown in the accompanying figures. A protective atmosphere of forming gas (10 pct H2, 90 pct N2) was continuously passed through the gas-tight annular volume between the core and outer jacket. The water-cooled copper ends shown in Figs. 1 to 3 sealed the ends of the tubes. Annealed molybdenum wire (40 mil) was wound on the recrys-tallized alumina core (35.5 mm OD by 28.5 mm ID by 30 in.) at a spacing of 0.36 in. The spacing was reduced to 0.29 in. close to the ends to partially compensate for end heat losses. Fig. 2 shows two molybdenum-ground lead taps off the center of the winding to produce the two independect temperature zones. The ground lead was insulated with 1/4-in. alumina tubing and connected to the grounded copper end by passing the wires through the gas port tube and crimping each wire into notches cut at the mouth. To prevent sagging of the core at high temperatures, refractory "troughs" were cut from alumina and mullite tubing to cradle the core. The support "troughs" which were 6 in. long and located about 5 in. from each end of the annular volume also served as radiation shields. In view of the different expansion coefficients and operating temperatures of the core and outer jacket (3.5 in. OD by 3 in. ID by 30 in.), the ends of the tubes were smoothed and squared to produce good sealing surfaces, and the core was ground 1/32 in. shorter than the outer tube. The power terminals were passed through their respective conductor ports which were internally lined with alumina tubing and sealed with Armstrong adhesive A-6 leaving about 1 in. of protruding lead wire for external connections. (Adhesive A-6 is a two-component system consisting of an epoxy-type resin with a mineral fibrous filler and an amine catalyst.) Figs. 1 and 3 show the entire assembly of concentric refractory tubes, copper ends, and sealing o rings, held together with two brass flanges and 3/8-in. bolts with heavy springs for dimensional flexibility. The furnace shell (19 by 19 by 30 in.) consisted of a steel frame with sheets of transite forming a box. The copper ends protruded through two 3.75-in. circular holes cut at the center of each transite square end. Thermal insulation next to the transite exterior was effected by K-23 B&W bricks alld in a central volume of approximately 9 by 9 by 30 in. a row of K-30 bricks was lined up to support the hot assembly. The rest of the volume around the outer

Jan 1, 1965

By AIME AIME

AT A meeting of the Boston Local Section of the Institute, on Oct. 3, approval was voted to the work done by its representatives on the Committee opposing the passage of a bill by the, Massachusetts Legislature to license engineers. The report is as follows: The Joint Committee on Licensing of Engineers, which included a Committee of the Boston Society of Civil Engineers and representatives of seven other societies,* submitted to the members of the respective societies, on Jan. 8, 1921, a report in which it was stated that on Jan. 4, 1921, a petition was filed in the Legislature of Massachusetts by the Boston Chapter of the American Association of Engineers accompanied by a bill to provide for the registration of professional engineers and land surveyors. This report also stated that at a meeting of this committee, held on Jan. 7, 1921, it was voted:

Jan 1, 1921

By H. C. Fiedler

SILICON-IRON strip with a cube-on-edge secondary recrystallization texture is made commercially as thin as 10 mils. With inclusions present to inhibit normal grain growth, a few grains, and these have the (110)[001] orientation, grow very large at the expense of the matrix grains. In the absence of inclusions, only normal grain growth occurs and the texture is not as strong as when it is developed by secondary recrystallization.1 Strip for use at frequencies of 400 cps and higher is made in thicknesses of 1, 2, and 4 mils by a process developed by Littmann.2 Attempts to develop a strong texture in these thin gages by the same methods as were used for the 10 mil and thicker strip were unsuccessful,2 and Littmann's solution to this problem was to cold roll the grain

Jan 1, 1963

By A. E. Back, K. E. Tame, W. F. Wyman, S. F. Ravitz

IN 1940, when it appeared that the United States soon might be cut off from foreign sources of high-grade manganese ore, the Bureau of Mines began an extensive series of investigations on the production of ferrgrade products from low-grade domestic ores by ore-dressing, pyrometallurgical, and hydrometallurgical methods. Among the hydrometallurgical processes for oxide ores investigated was a sulphur dioxide process in which the manganese was recovered from solution by precipitation with lime;1 this process is essentially as follows: The ground ore is suspended in water and leached with SO2 gas to extract the manganese as manganese sulphate. Calcium chloride is then added to the slurry to precipitate calcium sulphate and form manganous chloride solution, and the resultant slurry is filtered to remove the calcium sulphate together with the in- soluble portion of the ore. Calcium hydroxide is added to the filtrate and the mixture is agitated with air to precipitate hydrated oxides of manganese and regenerate calcium chloride. The precipitate is calcined, and the calcium chloride is recycled. To determine manganese extraction, reagent consumption, grade of product, and similar data for this process, a laboratory investigation was started. Two minor modifications in the process were soon made: (I) the ore was suspended in calcium chloride solution instead of in water in order to extract the manganese and precipitate the calcium sulphate in one step, and (2) the agitation with air after the addition of slaked lime to the filtrate was omitted because manganese hydroxide is readily precipitated without air agitation and can be nodulized or sintered to a satisfactory product. When the process was carried through a series of cyclic tests, it was noted that the concentration of dithionate ion (which is one of the products of oxidation of SO2 by MnOz) in the cycled solution gradually increased; at the same time the chloride content of the solution gradually decreased, owing to soluble losses in the leach residue and in the precipitate. After a number of cycles, the calcium chloride became virtually completely replaced by calcium dithionate, the quantity of dithionate ion formed during the leach being more than enough to make up for the losses of dithionate in filtering. The calcium chloride process thus developed into a calcium dithionate process. Since the latter eliminated the need for make-up calcium chloride, and the dithionate solutions appeared much less corrosive than the chloride solutions, work on the calcium chloride process was dropped in favor of the calcium dithionate process. ACKNOWLEDGMENTS The dithionate process was developed by the Bureau of Mines as part of a

Jan 1, 1946

By M. M. Wong, D. E. Couch, G. M. Martinez

The Bureau of Mines electrowon hafnium metal with an average oxygen content of' 150 ppm at 700°C from an electrolyte containing 27 wt pct LiCl, 62 wt pct RbCl, and 11 wt pct HfC14. The average anode and cathode current efficiencies were 90 pct at anode and initial cathode current densities of 86 amp per sq ft. Haf-nium metal with an average oxygen content of 440 ppm was electrowon at 800oC from an electrolyte containing 90 wt pct KC1 and 10 wt pct HfCl4. The average anode and cathode current efficiencies were similar to those obtained in the LiCL-RbCl-HfCl, electrolyte. The chlorine gas given off at the graphite anode was vented through either a silica or a graphite tube to prevent cell corrosion. THE current method for the commercial production of high-purity hafnium is the thermal decomposition of Hfl4.1 The iodide method is not adaptable to continuous process techniques. Nettle, Hiegel, and Baker2 studied the electrorefining of hafnium from hafnium sponge containing 800 ppm oxygen. They failed to obtain hafnium with 600 ppm oxygen in their initial deposits and obtained AEC specification for oxygen only after 75 pct of the soluble hafnium had been removed from the electrolyte. Calculations using their data indicated this was approximately 4 lb of hafnium. The electrolyte was then used to produce approximately 3 lb of hafnium with a low oxygen content. However, no data are shown concerning the amount of anode material initially used or what percent of it was dissolved, therefore, results are not suitable for evaluation of a continuous operation. In general, it was not possible to consistently obtain low oxygen content metal with the electrolytes described by Nettle, Hiegel, and Baker. Wong, Hiegel, and Martinez3 investigated the electrorefining process for hafnium and showed that even by strict control of electrolyte composition only relatively low oxygen reduction could be obtained. The oxygen contained in the hafnium anode material tended to transfer to the cathode deposit and only a limited purification was possible. Both the "iodide" and the "electrorefining" processes depend upon hafnium sponge as a starting material. The sponge is normally produced by magnesium reduction of HfC14 ' and does not meet AEC specifications for hafnium metal. Since only 30 pct of the anode feed could be utilized3 in the electrorefining cells, the Bureau of Mines developed an electrowinning process. HfC14 was used as the feed material for the electro-winning process described in this report. Many of the electrolytes used in the electrorefining studies3 ap- peared to be suitable carrier-electrolytes for HfC14. However, in the initial studies on electrowinning, it was desirable to use electrolytes that had low solidus temperatures and could be operated over a wide temperature range to investigate parameters of the process. Therefore, electrolytes containing LiC1, NaC1, KC1, RbC1, CsC1, and HfC14, in various combinations were explored. EQUIPMENT Chlorinator. Hafnium carbide was chlorinated to produce HfC14 in the batch-type chlorination shown in Fig. 1. Chlorination temperatures were measured with a thermocouple placed in the center of the HfC charge. A flow meter was used to monitor the helium and chlorine. The exhaust side of the silica chlorina-tor tube was equipped with a flask for collecting organic material released during the initial heating of the HfC. The temperature of an internal heater, which extended from the HfC14 condensing flask to the hot end of the chlorinator, was adjusted to prevent the HfC14 from condensing before entering the collection flask. Helium and excess chlorine were exhausted through the lid of the collection flask to an aqueous NaOH solution. Sublimer. Initial studies were conducted using a sublimer, Fig. 2, made by placing a 13/8-in. OD nickel thimble 11 in. long, inside a 11/2-in. ID nickel bell 12 in. long, and locking it in place. This unit was loaded with HfC14 and partially immersed in the molten electrolyte for sublimation directly into the electrolyte. In another sublimer shown in Fig. 3, the HfC14 was contained in a "resin reaction flask". Quartz wool, previously heated to 600aC, secured between two nickel wire screens, was placed just above the HfC14 powder. The lid contained a vacuum outlet, a gage, an argon inlet, and an air-cooled pipe for condensing the HfC14. This sublimer was evacuated and heated. The sublimation temperature was not critical and the sublimer operated satisfactorily at all temperatures between 250" and 350°C. Electrolytic Cell. The electrolyte chamber, Fig. 4, was made of mild steel 8-in. schedule 20 pipe, 30 in. long. The exterior was metallized with a Ni-Cr alloy. The electrolytes were contained in a 16 gage nickel or iron liner with a nickel heat shield on top. The cell was heated by a resistance furnace. A 21/2-in. ID by 25 in. long air lock was connected to one port of a two-port cell cover assembly through a slide valve. The cover assembly of the air lock was electrically insulated from the cell and was equipped with a rubber sleeve that provided for the passage of the cathode lead. This allowed the cathode deposits to be removed and a new nickel cathode to be introduced without allowing air to enter the cell. A tube-rod assembly was bolted to the other port on the cell cover assembly and was sealed by a packing seal. The tube-rod assembly consists of a graphite

Jan 1, 1970

By Bernard Kopelman

The particle size of tungsten metal powder used to make tungsten wire for use in radio tubes and incandescent lamps must be closely controlled if the highly desirable feature of nonsagging is to be achieved. Our knowledge of factors affecting the particle size of tungsten powder has been limited by lack of any extensive literature on the subject. There are no publications in the English language discussing this problem in any but the most general terms. In this paper the published experimental information on this subject is reviewed. Based on the published data and some experimental data obtained by the author, a general theory of factors affecting the particle size of tungsten metal powder obtained by reduction of the oxides in hydrogen is offered. The various forms of the tungsten oxides and tungsten metal used in the experiments described were those prepared by commercial processes, so that the experimental results obtained on a laboratory scale would be more directly applicable to actual commercial reduction conditions. The effects of time of reduction, temperature. and concentration of water vapor in the hydrogen were observed by counting microscopically the particle distribution obtained. Such a method is very time-consuming, but the data obtained by this method are more reliable and more fundamental than those obtained by any of the several instruments available for measurement of particle size. Review of the Literature Until the year 1912, tungsten powder was prepared commercially by reducing tungstic oxide in the presence of carbon or one of its compounds (such as glycerin or ethylene glycol), usually by firing the mixture in a closed fire-clay crucible. The resultant powder contained about 95 per cent tungsten.' When reduction of tungsten oxides by H² was introduced Commercially,²,³ several investigators became concerned with the equilibria between tungsten metal and its several oxides in an atmosphere of hydrogen and water vapor.4-l3 While these many investigations were in fair agreement on the equilibria WO³ + H2 = WO2 + H2O and WO2 + 2H2 = W + 2H2O there arose considerable controversy as to the chemical composition of the oxide between WO³ and WO², because of very poor agreement on the equilibrium constant of this intermediate oxide reduction reaction. The controversy on this intermediate oxide has been in progress for more than 100 years. Thus Malguti,14 in 1835, decided it was WO2.56 In 1857, Riche15 proposed the formula W³O8. Usler,16 in 1865, confirmed the existence of WsO8 and further proposed W1O1l. In 1897, Desil7 agreed upon W²O8 but also found W5O14.

Jan 1, 1947

By Bernard Kopelman

The particle size of tungsten metal powder used to make tungsten wire for use in radio tubes and incandescent lamps must be closely controlled if the highly desirable feature of nonsagging is to be achieved. Our knowledge of factors affecting the particle size of tungsten powder has been limited by lack of any extensive literature on the subject. There are no publications in the English language discussing this problem in any but the most general terms. In this paper the published experimental information on this subject is reviewed. Based on the published data and some experimental data obtained by the author, a general theory of factors affecting the particle size of tungsten metal powder obtained by reduction of the oxides in hydrogen is offered. The various forms of the tungsten oxides and tungsten metal used in the experiments described were those prepared by commercial processes, so that the experimental results obtained on a laboratory scale would be more directly applicable to actual commercial reduction conditions. The effects of time of reduction, temperature. and concentration of water vapor in the hydrogen were observed by counting microscopically the particle distribution obtained. Such a method is very time-consuming, but the data obtained by this method are more reliable and more fundamental than those obtained by any of the several instruments available for measurement of particle size. Review of the Literature Until the year 1912, tungsten powder was prepared commercially by reducing tungstic oxide in the presence of carbon or one of its compounds (such as glycerin or ethylene glycol), usually by firing the mixture in a closed fire-clay crucible. The resultant powder contained about 95 per cent tungsten.' When reduction of tungsten oxides by H² was introduced Commercially,²,³ several investigators became concerned with the equilibria between tungsten metal and its several oxides in an atmosphere of hydrogen and water vapor.4-l3 While these many investigations were in fair agreement on the equilibria WO³ + H2 = WO2 + H2O and WO2 + 2H2 = W + 2H2O there arose considerable controversy as to the chemical composition of the oxide between WO³ and WO², because of very poor agreement on the equilibrium constant of this intermediate oxide reduction reaction. The controversy on this intermediate oxide has been in progress for more than 100 years. Thus Malguti,14 in 1835, decided it was WO2.56 In 1857, Riche15 proposed the formula W³O8. Usler,16 in 1865, confirmed the existence of WsO8 and further proposed W1O1l. In 1897, Desil7 agreed upon W²O8 but also found W5O14.

Jan 1, 1947

By Willard R. Jillson

Within the last ten years Kentucky has become celebrated for its low-sulfur bituminous coals. Prior to this time, many investigators had discnvered the abundance of this coal but the fact was unknown to the general public until the last extension of Kentucky's mountain railroads' was completed. The coals of. Kentucky are, broadly, separated into two geographic units—the eastern and the western coal fields. On a basis of their sulfur content these coal fields fall into three units, which are indicated on the accompanying map, by numbers, 1, 2, and 3. The area marked 1, the very southeastern part of the state, shows the lowest sulfur, all coals being averaged. In this field the maximum is 1.04 per cent. and the minimum as low as 0.68 per cent.—both averages and not individual analyses, which would of course show much greater variation. Included in this group are the following counties: Lawrence, 0.87 per cent.; Martin, 0.75 per cent.; Johnson, 0.73 per cent.; Magoffin, 0.87 per cent.; Floyd, 0.88 per cent.; Pike, 0.68 per cent.; Knott, 1.04 per cent.; Perry, 0.77 per cent.; Letcher, 0.80 per cent.; Leslie, 0.70 per cent.; Harlan, 0.79 per cent.; Knox, 0.86 per cent.; and Bell, 0.92 per cent. The total average sulfur content for these thirteen counties is 0.82 per cent. Bordering the restricted low-sulfur field of eastern Kentucky is the No. 2 belt (slightly darker), 'which is commonly known as the western border of the eastern coal field. Here the maximum sulfur is 2.91 per cent. and the minimum 1.05 per cent. Again these are averages and not individual analyses. The counties in this group are: Greenup, 2.60 per • cent.; Boyd, 1.67 per cent.; Carter, 1.07 per cent.; Morgan, 1.40 per cent.; Wolfe, 1.83 per cent.; Lee, 2.93 per cent.; Breathitt, 1.85 per cent.; Owsley, 1.09 per cent.; Jackson, 1.06 per cent.; Rockcastle, 2.30 per cent.; Clay, 1.09 per cent.; Laurel, 1.08 per cent.; Pulaski, 2.24 per cent.; Whitley, 1.05 per cent.; McCreary, 1.50 per cent.; and Wayne. The sulfur percentage average for this area including fifteen important counties is 1.65 per cent. In the western Kentucky field, the great increase in the percentage of sulfur is at once apparent. Ohio County with 3.78 per cent. is the highest and Hancock with 2.87 per cent. is the lowest. The percentages by counties is as follows: Hancock, 2.87 per cent.; Henderson, 3.22 per cent.; McLean, 3.27 per cent.; Muhlenberg, 3.31 per cent.; Hopkins,

Jan 1, 1920

By Thomas S. West

A gravel pack completion is described which, in addition to sand screening, permits the exclusion of water and gas from the oil production. A gravel pack composed of alternate layers of different particle sizes is employed which has horizontal but not overall vertical permeability to drilling mud. A packer is run on the tubing and set in the screen member, around which the gravel layers are placed. The hydrostatic pressure of a mud column prevents the entry of fluid above (or below) the level of the packer. The producing level may be changed by raising or lowering the tubing. Vertical coning is controlled by replacing the mud column with oil and continually injecting oil into the sand section above (or below) the packer. Three well histories with this completion are described in detail. Statistics on the first 100 completions of this general type are given. INTRODUCTION Gravel pack type completions have heretofore been almost exclusively for the purpose of preventing sand production. This paper describes and presents well data on a gravel pack procedure which is primarily a selective completion method. Although complete sand screening is achieved, the exclusion of gas and water is the primary objective. The application of this procedure is therefore not limited to unconsolidated sands. Application to hard sand or limestone reservoirs is equally feasible. GRAVEL PACK PROCEDURE Figs. 1 and 2 illustrate the gravel pack procedure employed. The gravel is placed by means of the frangible disc type dump bailer shown by Fig. 1.' This bailer has its lower end closed by a plate glass or steel rimmed cement disc. This disc breaks when the bailer passes over the liner allowing the gravel to be deposited around the liner. The liner employed is usually 2 3/8 in. to 2 7/8 in. OD and 1.995 in. or 2.07 in. ID (ID of 2-in. tubing or 2-in. line pipe). Two tvDes,. of slotted sections have been employed. One type employs horizontal hacksaw cut or milled slots of about 1/32 in. width opposite the producing section. The other type employs the same type,. of horizontal slot. but slots are placed in groups of six at intervals of about 8 in. A cage composed of half round vertical bars, spaced apart, is then welded to the pipe over the slotted sections. This liner has the bottom end closed, usually by bull plugging. The upper end is closed, while placing the gravel, by a cap which is held in place by means of a shear pin. This cap has a mandrel attached to it which extends downward inside of the liner to within a few inches of its lower end. The top of the liner cap is formed to fit a releasing type wire line core barrel overshot. The liner is lowered to bottom on the wire line by means of this overshot and released. The length of the liner is usually suffi- cient to extend to about 20 ft above the highest level to which gravel is to be placed, the upper end of the liner always being in the casing. The dimensions of the liner and bailer are such with respect to the casing size that the bailer will always pass over the liner. After releasing the liner the bailer is run to bottom. Passage of the empty

Jan 1, 1951

By Thomas S. West

A gravel pack completion is described which, in addition to sand screening, permits the exclusion of water and gas from the oil production. A gravel pack composed of alternate layers of different particle sizes is employed which has horizontal but not overall vertical permeability to drilling mud. A packer is run on the tubing and set in the screen member, around which the gravel layers are placed. The hydrostatic pressure of a mud column prevents the entry of fluid above (or below) the level of the packer. The producing level may be changed by raising or lowering the tubing. Vertical coning is controlled by replacing the mud column with oil and continually injecting oil into the sand section above (or below) the packer. Three well histories with this completion are described in detail. Statistics on the first 100 completions of this general type are given. INTRODUCTION Gravel pack type completions have heretofore been almost exclusively for the purpose of preventing sand production. This paper describes and presents well data on a gravel pack procedure which is primarily a selective completion method. Although complete sand screening is achieved, the exclusion of gas and water is the primary objective. The application of this procedure is therefore not limited to unconsolidated sands. Application to hard sand or limestone reservoirs is equally feasible. GRAVEL PACK PROCEDURE Figs. 1 and 2 illustrate the gravel pack procedure employed. The gravel is placed by means of the frangible disc type dump bailer shown by Fig. 1.' This bailer has its lower end closed by a plate glass or steel rimmed cement disc. This disc breaks when the bailer passes over the liner allowing the gravel to be deposited around the liner. The liner employed is usually 2 3/8 in. to 2 7/8 in. OD and 1.995 in. or 2.07 in. ID (ID of 2-in. tubing or 2-in. line pipe). Two tvDes,. of slotted sections have been employed. One type employs horizontal hacksaw cut or milled slots of about 1/32 in. width opposite the producing section. The other type employs the same type,. of horizontal slot. but slots are placed in groups of six at intervals of about 8 in. A cage composed of half round vertical bars, spaced apart, is then welded to the pipe over the slotted sections. This liner has the bottom end closed, usually by bull plugging. The upper end is closed, while placing the gravel, by a cap which is held in place by means of a shear pin. This cap has a mandrel attached to it which extends downward inside of the liner to within a few inches of its lower end. The top of the liner cap is formed to fit a releasing type wire line core barrel overshot. The liner is lowered to bottom on the wire line by means of this overshot and released. The length of the liner is usually suffi- cient to extend to about 20 ft above the highest level to which gravel is to be placed, the upper end of the liner always being in the casing. The dimensions of the liner and bailer are such with respect to the casing size that the bailer will always pass over the liner. After releasing the liner the bailer is run to bottom. Passage of the empty

Jan 1, 1951

By David Kellogg

THE electrodeposition of iron has been practiced for many years. The earlier work along this line was directed toward the preparation of pure metal, but later applications were the production of thin adherent coatings as used in the electrotyping of Russian banknotes in 1860. A historical discussion of this subject, however, would require too much space and would be inconclusive. Much has been written on the subject but no two experimenters use the same methods or materials. During the war, the British army repair, shops developed a method for building up worn parts of automotive machinery, aero engines, etc., using the cold surfate . bath and low-current density method.1 The method has been used successfully in commercial work for the production of about 6000 repaired parts, therefore experiments along the same lines have since been made at the Westinghouse research laboratory. The apparatus used was essentially that described in the article mentioned except that the current -used for cleaning was obtained from a 3/4-kw. 60-volt, house-lighting generator, direct connected to a 2-h.p., 870-r.p.m., 440-volt, three-phase, type CS induction motor; while the plating current was taken from the storage batteries at any desired voltage. The anodes were made from 0.036 in. (0.9 mm.) Armco iron made into cylinders 5 in. (127 mm.) long by 8 in. (203 mm.) in diameter with a disk of 1/8 -in. (3.17 mm.) micarta, having a 2 in. hole in the center fitted at each end of the cylinder for stirring. The anodes were hung on a wooden rocker frame driven by a wooden connecting I rod directly connected to a small reducing gear, such as is sold by most of the apparatus supply firms; 3 gal. stoneware crocks were used as containers. The solution used by Thomas, 75 gm. of the crystallized ferrous ammonium sulfate per liter, was tried, using the current density recommended by him; namely, 0.33 amp. per sq. dm. Under these conditions the deposit on a ½ -in.

Jan 2, 1922

By C. E. Ells, A. Sawatzky

The effect of neutron irradiation on tensile properties of the Zr-2.5 wt pet Nb-0.5 wt pet Cu alloy has been evaluated for integrated neutron fluxes up to 3 x 1020 n per sq cm (E > 1 Mev). Specimen temperature during irradiation was normally 300°C. The material was studied both in annealed and in quenched and aged conditions. When the alloy (fab-ricatedfrom sponge zirconium) is quenched from a temperature =40°C below the a + ß/ß transformation and aged 6 hr at 535oC, neutron irradiation up to the maximum studied has little effect 071 reduction in area, although there is a drop in uniform elongation and the yield strength is increased by = 25 pct. Conversely, irradiation of the alloy in an annealed condition can result in much greater changes in properties, with yield-strength increases of up to 200 pet. The dose dependence of irradiation hardening obeys the saturation equntion proposed by Makin and Minter to explain hardening in copper and nickel: Act = C[1 - exp(-Døt)/1/2 A rnodel is proposed to explain the effect of niobium content and metallurgical condition on irradiatiotz behavior. THE Zr-Nb alloys form a class of material susceptible to marked strengthening by quench and age heat treatment. With niobium concentrations as low as 2.5 wt pet, strengths can be developed which are double those of annealed zircaloy-2.1 Since the neutron-capture cross section of the Zr-2.5 wt pet Nb alloy is nearly identical to that of Zircaloy-2, significant gains in power-reactor neutron economy could be obtained by replacing stressed in-reactor components of Zircaloy-2 with the heat-treated Zr-Nb alloy. When the solution heat-treatment temperature is in the high (a + ß) phase, then the tensile properties of the quenched (and aged at 500°C) material have been shown to be relatively insensitive to neutron irradiation.1-3 Addition of small quantities of copper to the binary Zr-2.5 wt pet Nb alloy gives a useful reduction to the corrosion rate in air and carbon di- oxide,&apos; and a ternary alloy of composition Zr-2.5 wt pet Nb-0.5 wt pet Cu was chosen for one component in the Douglas Point Reactor. This application utilizes both the corrosion resistance to a moist air-carbon dioxide environment and the strength developed by a quench and age heat treatment. By suitable choice of solution heat treatment and aging temperatures, tensile properties of the ternary alloy can be made nearly identical to those considered as optimum for the binary alloy. One effect of the copper addition, however, is to give marked changes in the aging kinetics of the alloy.&apos; For this reason, it appeared that neutron irradiation might promote overaging in the ternary alloy, particularly at the elevated service temperature (=300°C). This paper describes the effect of neutron irradiation on the tensile behavior of the Zr-2.5 wt pet Nb-0.5 wt pet Cu alloy in several metallurgical conditions pertinent to its use as a reactor material. The metallurgical conditions included those for which aging occurs in the unirradiated material, and fully annealed conditions for which a higher degree of thermal stability could be expected. 1) EXPERIMENTAL Two batches of alloy were used, both made from sponge zirconium. The fabricator&apos;s analysis of principle constituents showed no significant difference between the two batches, Table I. The rod material was rolled to 1.5 in. diam, and then swaged to final sizes of 0.5 to 0.375 in. diam at nominal temperatures of 785" and 625°C for the AM and AS batches, respectively. Subsequent hydrogen analyses at CRNL confirmed that the hydrogen concentration of the as-received rod was indeed <20 ppm. However, it was found that the (a + ß)/ß transus for the AM material was approximately 25°C lower than that of the AS material, indicating rather more difference in oxygen concentration than given in Table I. Except for a small difference in yield-point behavior, the tensile prop-

Jan 1, 1965

By Orson Shepard

TEXTBOOKS on fire assaying list the following standard methods of assaying sulphide ores: (1) scorification, (2) litharge-niter, (3) soda-iron, (4) roasting, (5) combination wet and fire. The litharge-niter method has proved the most satisfactory, and is the usual method for assaying sulphide ores and concentrates. It gives good results when suitably conducted and when lead buttons of uniform size are produced so that the cupellation temperature can be properly controlled. The main objection to the method is the inconvenience of determining the amount of niter to add to each assay charge to produce lead buttons of suitable size from a set of fusions. For the assay of different materials, the niter requirement varies with the particular sulphide minerals present, the amount of sulphide minerals present, the slag composition, and the rate of heating the charge in the furnace. Except for some materials, so uniform in composition that the same amount of niter can be used in assaying all samples, the amount of niter must be determined for each sample to be assayed. This is done by either: (1) calculation from an estimate of the kind and amount of sul-phide minerals present in the sample, or (2) a preliminary fusion. Some assayers use one method and some the other. The estimation method takes the least time; but much experience is required to obtain close estimates on a wide variety of materials. At best, the sizes of the lead buttons obtained vary considerably. The preliminary fusion method usually gives lead buttons of fairly uniform size, although sometimes there is a variation because the oxidizing power of niter varies a little with the charge. This variation can be eliminated by two preliminary fusions: one to determine the reducing power of the ore, and the other to determine the oxidizing power of niter in the charge. The objection to the prelimi-nary fusion is the extra time and expense involved.

Jan 1, 1939

By E. T. Turkdogan, J. H. Swisher

The solubility of oxygen in zone-refined iron was determined in the temperature range.from 881" to 1350°C. The solubility in a iron at 881°C ms found to be about 2 to 3 ppm; in y iron, the solubility was found to increase from about 2 to 3 ppm at 950°C to about 25 ppm at 1350°C. The permeability of oxygen in an iron -0.1 pct Al alloy was determined in the y iron range, using an internal oxidation technique. By combining the permeability and solubility data, the diffzisivity of oxygen in y and a iron was calculated. The oxygen diffusivity in solid iron may be s~immarized as follows: 8820 For 6 iron and IN a recent paper, Hepworth, Smith, and Turkdogan1 reported on the solubility, permeability, and diffusivity of oxygen in 6 iron, together with permeability data for oxygen in a iron. In the present investigation, similar measurements were made in the ? phase region. In addition, a solubility measurement was performed in the a phase region to permit calculation of the diffusivity of oxygen in a, iron. References to early work on this subject are given in the previous publication.&apos; EXPERIMENTAL Solubility Measurements. The oxygen solubility was determined by equilibrating cylindrical samples 0.3 in. diam by 14 in. long of zone-refined iron in water vapor-hydrogen gas mixtures. The zone-refined iron was prepared by B. F. Oliver of this laboratory, using an apparatus and technique described elsewhere.2&apos;3 Six zone-melting passes were used to achieve a total impurity level of about 30 ppm. Of this 30 ppm, the combined nickel and cobalt content was about 20 ppm, oxygen was 4 ppm, and all oxidizable impurities less than 1 pprn each. A vertical resistance furnace wound with molybdenum wire was used for the experiments. The temperature was measured before and after each experiment with a Pt/Pt-10 pct Rh thermocouple. In the gas train, flow rates of hydrogen and argon were measured with capillary flow meters, and the resulting mixtures were passed through a column containing 90 pct oxalic acid dihydrate and 10 pct anhydrous oxalic acid to obtain predetermined ratios of H2O to H2. The vapor pressure of H2O above this mixture as a function of temperature is well-known.4 The exit gas was ana- lyzed periodically for H2O, and good agreement (+3 pct) with the calculated composition was obtained. The zone-refined iron specimens were held in the furnace for a sufficient length of time for equilibration, e.g., 18 hr at 1350°C and 1 week at 881°C, then quenched in a brine solution. After removing the surface oxide from the samples by machining, duplicate analyses were obtained by a combined vacuum fusion-infrared method; oxygen analysis was reproducible within 2 ppm. Permeability Measurements. The permeability of oxygen in alloys containing about 0.1 pct A1 was determined by internal oxidation and measurement of the subscale thickness as a function of time. The experimental alloys were prepared by adding aluminum to electrolytic iron (grade 104A plastiron) that had previously been vacuum-carbon deoxidized. The resulting ingots contained about 20 ppm O, 100 pprn C, 40 pprn Si, 50 ppm Mo, 20 ppm P, 20 ppm S, and 20 pprn Zr as the principal impurities. After hot rolling the ingots to 1-in.-thick slabs, specimens were machined from the stock in the form of rectangular plates, 4 by 5 by 2 in. The general procedure for the permeability experiments was the same as those for the solubility measurements. The specimens were cooled in a reducing atmosphere rather than quenched, however, in order to maintain a clean surface for measurement of the sub-scale thickness. This measurement was made on a polished cross section of each specimen, using a microscope with a micrometer stage. The inclusions formed at the lowest temperature 1033°C were too small to be seen with an optical microscope. An electron micrograph showing the size and shape of individual particles is shown in Fig. 1. The dark band in the picture is a boundary between two subgrains. The subscale thickness in these samples was measured with an optical microscope after heavily etching in 2 pct nital, which gave contrast between the subscale and the unoxidized zone.

Jan 1, 1968

By J. H. Bowden

The water-hoisting tanks herewith illustrated have been designed for removing large quantities of water from recently-flooded mines, through their hoisting-slopes, with rapidity proportional to the capabilities of the hoisting machinery available, the tanks being adaptable to slopes of small sectional areas and varying pitches. The following features may be of interest to those operating mines liable to be flooded. 1. The arrangement of doors on each tank by which it may be automatically filled by immersion at any point on the varying pitch of a slope, the water being retained while hoisting on the flat as well as on the steep pitches, and quickly and autoniatically discharged at the top. 2. The arrangement, in connection with the above-mentioned doors, of side-wheels over the rear wheeh of each tank and of sidedumping tracks at the top of the hoist, for the automatic emptying and quick return of the tank. 3. The arrangement of two or more tanks, one in front of the other, so as to give as large a capacity in a single hoist as the enginepower will permit, yet without making any tank too large for the sectional area of the slope, or of unwieldlp length, or of such a shape that it cannot easily pass over vertical curves; and without concentrating too much weight on any one pair of wheels or on any point of the track. The details of construction and method of operating are shown by the accompanying sketches (Figs. 1 and 2), and the following more detailed description. At the end of each tank is a large iron door of almost the full size of the end of the tank, opening inwards, so that when immersed the tanks fill almost instantly. To provide for holding the water while it is hoisted up flat pitches, a wooden door is attached to the front of each tank, opening outward. Each front door is attached to the

Jan 1, 1892

By Robert L. Gower, John H. Cash

Interconnections for integrated circuits operating at rnicrowaue frequencies rzust be formed as microwave transmission lines. This paper describes the fabrication of one type of microwave transmission line known as microstrip, which can be fabricated on momlithic integvated circuits by deposition of 0.1- to 2.0-mil-thick metal and dielectric films. A microstrip line consists of a narrow conducting strip separated from a gvound plane by a dielectric material. Thick films of good conducting materials are required for low loss at high frequencies. Vacuum-evaporation methods for depositing aluminum and silver films are described. The narrow conducting strip forming the top conductor is defined by photoresist-etching techniques after the thick rnetal film is deposited. Gold plating with a photoresist-masking technique is also used to forrrz thick metal conductors. The microstrip dielectric layer is formed by deposition of suitable dielectric films 0.1 to 2.0 mils thick. Several deposition techniques were investigated. Electron-beam evaporation of quartz is described. Electrical characteristics of deposited microstrip lines at microwave frequencies are included and the microwave power attenuation as a function of conductor and dielectric thickness is given. It is concluded that the deposition of microstrip transmission lines is a practical way to achieve low-loss interconnections in monolithic integrated circuits. CONNECTIONS between devices in low-frequency and digital circuits are usually made by connecting the devices with a length of conductor such as a metal lead oi- a wire. If this method were applied to microwave integrated circuits, the reactance of the connection due to the series inductance and parasitic capacitance would be very high, in fact too high to permit successful circuit operation. In addition, resonance and antenna effects would occur at high frequencies, especially when the connections approach the wavelength.&apos; Connections which are sections of transmis- sion line will overcome these difficulties. The electromagnetic field is confined to the volume of the transmission line, parasitic effects are not present, and circuit design utilizes familiar transmission-line techniques. TRANSMISSION LINES One type of transmission line geometrically adaptable to integrated circuits is known as micrstri. A microstrip line can be generated from a coaxial line as shown in Fig. l(a). The outer conductor of the coaxial line is distorted into an ellipse (b) and finally separated at the extremities of the ellipse to form the strip transmission line shown in (c). One of the ground planes is then removed, leaving a conducting ground plane and a narrow conducting strip separated by a dielectric. Two configurations of the microstrip line were considered in this investigation, Fig. 2. Type (a) is fabricated from a semiconductor slice with the ground plane and top conductor deposited on each side of the slice. The semiconductor slice serves as the die lec-

Jan 1, 1967

By B. Trafton, H. L. Hartman

An earlier paper1 introduced a rapid solution for mine ventilation network problems, employing the digital computer. In this paper, refinements and additions to the computer program which expand its scope of application and a solution to a ventilation network having a coal-mine configuration are presented. The revised program-written in FORTRAN for the IBM 7074 computer in Penn State&apos;s Computation Center-is applicable to any mine ventilation network employing natural splitting and having its fan located externally. It accomplishes the following: 1) Distributes an assumed total quantity of air flow among all airways of the network, performing iterations until an acceptable level of error is attained and the flow in each airway is known. 2) Computes the pressure drop (head loss) in all airways and the overall network and mine head losses. 3) Determines the operating point for the mine and a given fan, adjusting all air flows and head losses to the actual values. The basis of analysis used is the Hardy Cross method of successive approximation for solution of any fluid network, explained in detail in the earlier article. In brief, corrections to assumed air flows are computed for all the loops of a network; following distribution to the airways of each loop, the process is repeated until acceptable accuracy is attained. When a hand or desk-calculator solution is used, there is advantage to close approximation of the actual flows in the original assumption of air-flow distribution in the network. Since publication of the previous paper, in which assumption of initial quantities was also necessary in the computer solution, the program has been modified to make the first distribution itself, assigning the total quantity to one specified path in a split and then proceeding to correct the flows. The number of iterations required for a high-accuracy solution is little affected, at least in networks of few loops, 10 to 15 being sufficient to reduce the error to 1 pct or less. The principal limitation of the original program was that it contained no provision for handling a fan. Further, it did not determine the overall head loss in the network and was not designed for multiple parallel openings, such as occur in coal mines. The revised program was written to overcome these disadvantages and also to print out the adjusted flows for all airways, corrected for the system operating point. At the present time, the program can handle only an external fan. The procedure employed in solving a ventilation network is now as follows: 1) Input data consisting of airway characteristics and loop designation along with assumed direction of flow, for each branch of the network and any trunk airways, are fed into the computer. 2) Employing the resistance relationship, where R is resistance, K is friction factor, P is perimeter, L is length, A is area, and n is number of airways (where multiple intakes and returns are employed); values are computed for all airways. 3) An arbitrary, continuous route of flow through the circuit—from beginning to end—is selected and the path identified by airway numbers and read in. 4) Based on a total assumed quantity of flow Q (usually 100,000 cfm.) the computer employs the procedure described in the previous article in distributing quantities of flow among all the airways until a satisfactory level of accuracy has been attained (in the current program, the machine is turned off when the cumulative corrective error in cfm, is reduced to five times the number of airways). 5) Airway head losses H are calculated by the relation H = R Q2 and cumulated along the original designated flow route to find the overall head loss for the network. Trunk airways are then included to obtain the mine head.

Jan 1, 1964