Search Documents

By E. S. Leaver

THE presence of soluble base metals in precious-metal ores usually precludes cyanidation as the best method of treatment. The laboratory experiments described in this paper show the possibility of cyanidation of gold-silver ores containing less than 0.5 per cent, of cyanide-soluble copper and the effect of zinc in the cyanide solution. PART I.-COPPER IN CYANIDATION The proposed process of cyanidation of precious-metal ores containing copper is based on the regeneration of about 80 per cent., of the cyanide used in the dissolution of the copper. The method of regeneration involves a combination of the sulfide and the acid precipitation of the copper and silver. The resultant solution is made alkaline with lime, and the gold is precipitated with zinc dust. Solubility of Copper Minerals in Cyanide Solution In order to determine the solubility of copper minerals, including metallic copper, under the usual conditions for cyanidation, experiments were conducted with synthetic ores. Each ore sample was prepared by mixing the particular pure copper mineral with barren sea sand. Both the mineral and sand were ground to pass 100 mesh before they were thoroughly mixed. The percentage of copper in each head sample was determined by direct assay, excepting the metallic copper sample, which was computed on weights before mixing. Table 1 gives the results.

Jan 1, 1929

By John D. Verhoeven

A technique has been developed for carrying out normal freezing experiments with a current density of 2000 amp per sq cut passing through the solid-liquid interface. The equation relating the effective distribution coefficient to the equilibrium distribution coefficient in electric field-aided solidification, originally developed by Huckc et al.,1 has been modified for the case of concentrated solutions. Preliminary experiments on the Sn-Bi system give qualitative agreement with the equation. The data are analyzed by a slightly novel use of the normal freeze equation which allows one to determine the effective distribution coefficient more easily. Very extensive mixing in the liquid was found at these high current densities and it is postulated that the mixing results from a vertical component of the magnetic Lorentz force generated by the electric current. In the search for techniques of obtaining ultrahigh-purity metals the inefficient but very effective technique of electrotransport has received little attention. Electrotransport is most effective in the liquid state and a natural application, therefore, is to apply an electric field across the liquid zone of a zone-melting experiment. The present investigation was undertaken to study the effect of an electric field upon solidification of metals, so that the usefulness of electrotransport in such solidification experiments as zone melting could be determined. In zone-melting and normal-freezing experiments it is difficult to achieve complete mixing in the liquid in the immediate vicinity of the solidifying interface. Consequently a solute build-up will occur at the interface in the portion of the liquid where complete mixing does not occur (an equilibrium distribution coefficient, ko, less than one, and unidirectional atom motion will be implied throughout). This local solute build-up produces a corresponding rise in the solute concentration in the solid so that the ratio of the solute concentration between the solid and the bulk liquid is larger than the equilibrium distribution coefficient. This ratio is defined as the effective distribution coefficient, k,. The differential equation describing the solidification process may be derived by applying the continuity equation to an expression for the net solute flux at the interface. The solution to this differential equation then allows one to determine the solute distribution in the liquid and the relationship between k0 and ke. One of the most useful solutions to this equation was first derived by Burton, Prim, and Slichter,' in which they assumed that a) the equilibrium distribution of solute was maintained on the plane of the interface, 11) the solute build-up ahead of the interface in the liquid disappeared at a distance 6 from the interface, and c) the solute distribution in the liquid was invariant with time. The following well-known relation between ko and ke was then obtained, where R is the rate of solidification and D the diffusion coefficient of the solute in the liquid. This equation appears to correlate the data from a number of different types of solidification experiments very well. Application of an electric field across the solid-liquid interface can produce an additional flow of solute atoms as a result of the electrotransport. When the polarity of the field is such as to direct the electrotransport flux away from the interface the solute build-up may be diminished, even to the point of producing a solute depletion and a consequent ke smaller than ko. The quantitative description of this process and the resulting form of Eq. [I] was first given by Hucke et al.1* and then inde- where E is the electric-field intensity and U is the differential mobility, i.e., the velocity of the solute atoms with respect to the solvent atoms per unit electric field. Both authors follow the method of Burton, Prim, and Slichter in their derivation, the only difference being the additional electrotransport term in the flux equation. It has been pointed out1,3 that Eq. [2] predicts the possibility of a noticeable increase in the purification of materials by solidification in an electric field. The validity of Eq. 121 has not been checked experimentally and it is possible that other factors' arising from the presence of an electric field across

Jan 1, 1965

By S. L. Couling, G. W. Pearsall

MOTT and Hainesl have summarized the available techniques for the examination of aluminum and some anisotropic metals under polarized light. Many of these techniques are rather complicated and cumbersome. This note presents a simplified procedure for using polarized light to obtain information about the crystallographic orientation of any particular grain in a magnesium alloy. Since other hexagonal metals are more anisotropic than magnesium, similar methods might well be applied to them. Ernst and Laves2 produced a double-refracting layer on magnesium alloy specimens by SuCCeSSiVe etches in solutions of perchloric acid in alcohol, magnesium perchlorate in perchloric acid and alcohol, and calcium and magnesium carbonates in water. They examined their specimens under polarized light and were able to determine the direction of the hexagonal axis in any particular grain from a measurement of the extinction behavior of that grain. The disadvantages of their technique were the complexity of the etching procedure and the explosive hazard of the perchloric acid-alcohol solutions. Results equivalent to those of Ernst and Laves were obtained in this investigation after etching specimens for 6 to 8 sec in an acetic picral solution containing 100 parts 6 pct picral, 10 parts distilled water and. 5 parts acetic acid. A. similar etch has been used by George3 in earlier metallographic

Jan 1, 1958

By Richard L. Ash, J. H. Kelley, W. H. McCracken, William Bellano, J. D. Morgan, D. R. Mitchell

Rubber-tired shuttle cars are finding increased use as gathering units in under-ground mines. As the name implies, they shuttle back and forth between the loading machine and the section transportation system. In many mines their use has in-creased the efficiency of the loading and gathering operations, resulting in increased output per man employed and decreased costs per ton. A car motivated by cable reel is shown in Fig I, and a car with battery power in Fig 2. A brief chronological account of the development of rubber-tired haulage equip-ment follows: 1936. J. H. Fletcher1 installed his first unit for the Blue Bird Coal Co., Carrier Mills, Illinois. This unit consisted of a battery-powered industrial tractor and a trailer with drop-bottom doors. The coal was transported for short distances to a hopper for discharge into mine cars. 1938. A battery-powered self-contained shut-tle car with rubber-tired wheels was built by the Joy Manufacturing Co. Two units were installed for the Katherine Coal Co., Lumberport, West Virginia. The shuttle car consisted of a materials-receiving body with a conveyor in the bottom to move the coal forward during loading and to unload the coal at the discharge station. 1938. A cable-powered shuttle car with four-wheel drive and four-wheel steer was developed by A. L. Lee. Special wheels mounted on king pins on the car permitted all four wheels to be steered. The cable reel was driven by a torque motor to give the cable proper tension during the payout and uniform tension when rewound on the reel. 1940. A Joy shuttle car equipped with trolley poles and a cable reel was developed to operate on a double trolley-wire system.2 This unit was installed at a potash mine of the International Minerals and Chemicals Corporation at Carlsbad, New Mexico. It is believed to have been the first installation of its kind, and the first of shuttle cars for handling material other than coal. 1941. Truck haulage3 was introduced in the Tri State District by the Eagle-Picher Mining and Smelting Company. Storage battery electric trucks were rebuilt to handle a bottom-dump trailer. 1941—Joy shuttle cars were installed in a mine of the United States Gypsum Company, Oakfield, New York. 1943—Joy shuttle cars were installed in several bauxite mines.4,5 1945—Joy shuttle cars installed in a lime-stone mine. Because of the nature of bituminous coal beds trackless equipment has had its greatest application in bituminous coal mining. Usually a shuttle car is filled by a continuous type of loader. In Fig 3 is shown a continuous machine loading coal into a battery-powered shuttle car, 48-in. coal. However, intermittent types of load-

Jan 1, 1949

By E. J. Schmitz

This section was obtained last year during an examination of the iron-ore resources of the Rich Patch Mountain region and along Craig's creek valley. The geological members below the Oriskany and Lower Helderberg are exposed along the gap in a beautiful arch, while the higher formations appear in a more or less disturbed state on both sides of this arch, and are partly covered by dèbris. I have thought it of interest to the Institute to reproduce the drawing of this section at the gap which I made in my note-book at the time of my examination; those parts of the section, which lie at both sides of the arch, are only approximated. I give also a number of notes and measurements, referring to the Clinton formation and especially to the characteristics of the fossiliferous ore-seam opened in the various tunnels along the arch (see section). I. In Tunnel C: For the first 110 to 120 feet to second shute, we find the ore only here and there in small deposits, and interchanging with, or changing into, either a very siliceous limestone or a sandstone. About 10 feet beyond shute we have : Sandstone roof. Clay-slate, 211/2 inches Brown lean ore, 1 inch. Sandy clay, 3 to 31/4 inches. 1. Dark .brown ore, 3 inches. 2. Fossil-ore, changing into sandstone or very siliceous limestone, 2 inches. 3. Brown lean ore, 6 inches. Clay and slate below, 1 foot. 194 feet from entrance : Sandstone roof. Clay-slate, 4 inches. Clay 15 inches. Fossil-ore, 63 to 7 inches. Clay and ore, If inches. Lean brown ore, 2 inchea. Clay to bottom, 2 feet.

Jan 1, 1896

By Royal Preston Jarvis

minerals of different specific gravities. It is simple in construction, easily operated, capable of treating large quantities in a short time, and highly efficient under various conditions. The question, whether the material to be jigged has first been sized, determines the two principal methods of jigging. Jigging preceded by close sizing, generally known as the Continental or German system, involves a more or less elaborate series of screens or trommels, with attendant cost for installation, operation, and repairs. Jigging without sizing, known as the English system, is, according to Munroe,1 a development of the hand-jigging formerly employed in Cornwall . and introduced by English miners to this country." In its simplest form, the method consists in jigging an ore-mixture previously crushed to some maximum size (although, in some cases, even this preliminary is omitted) oil a relatively coarse sieve, and then jigging again on a filler sieve, the material passing through the first sieve and bedding. While many modifications have been necessary to adapt it for use in mills of large capacity, where hand-work was necessarily, replaced by machines, the principle remains the same; the fact that the English system has been successfully employed, both in this country and abroad, is well known; and arguments have been made for its efficiency and applicability in a wider sphere than it has occupied hitherto. II. REVIEW OF PREVIOUS INVESTIGATIONS. The fact that treating a mixture of minerals under jigging conditions increased the amount of mineral saved, or, as Professor Richards aptly terms it, 1 ?the extra jig-catch," has long been known. To account for this fact a number of theories have been proposed. The work of Rittinger2 in this field has, for many years, been a classic in the literature of ore-dressing. For the purposes of my present paper, however, the work of two American investigators, Prof. H. S. Munroe and Prof. R. H. Richards, is chiefly concerned. Professor Munroe has given the results of an elaborate series of experiments,3 and his deductions, based largely oil theoretical grounds, of this work. After reviewing briefly the two systems of jigging, followed by a discussion of Rittinger's formulas and

Sep 1, 1908

R. Culver, Colorado School of Mines I appreciated Professor Starfield's presentation, because his approach to blasting research reflects my feelings that a lot of practical work can be done toward making short-term advances that can be used by the blasting industry. We should not forget this in our attempt to obtain solutions to such idealized problems as a mathematical theory of cratering. The blasting industry would like very much to have the efficiency of the basic blasting processes improved from the range of 20 percent up into the range of 50 percent. And many of these problems should be approached and attacked in the laboratories. Blasting research at the Colorado School of Mines for the past two or three years has been aimed at improving the pre-splitting process, because we feel it will become increasingly important to the construction industry. We have attempted to develop a unified theory of how pre-splitting works; and the more we investigate, the more variables we find. We may never obtain an entirely satisfactory theory of pre-splitting, but we have developed a tentative explanation. One of our first conclusions was that we are making very poor use of the explosive in current pre-splitting practice. But to appreciate this we must first consider how pre-splitting works. A series of parallel boreholes are loaded with low density charges which are detonated simultaneously. (Delayed detonation has been tried, but most people agree that simultaneous detonation is best, and this is what we found.) Stress waves traveling out from each borehole interact with the stress waves from the two adjacent boreholes influencing the direction of the fractures that occur and propagate from each borehole. It has been suggested in private communication with Mr. Wilbur Duvall of the U. S. Bureau of Mines, that the fracture between the adjacent boreholes is a result of the net particle velocities above and below the pre-split plane which tend to set up a net tension along the pre-split plane. Regardless of what the exact mechanism is that causes the fracture to propagate from one borehole to the next, we know that it, in general, relies upon an interaction of the stress waves. However, if there is a large compressive stress normal to the pre-split place, it can practically eliminate any possibility of obtaining the desired pre-split by preventing the desired fractures from getting started. Aside from obtaining a fracture along the pre-split plane, the other major aim of pre-splitting is to minimize the fracturing that occurs around the borehole. We do this by drilling a large hole and putting a small charge in the middle of it. This results in a badly decoupled charge which is very inefficient; most of the energy is used in compressing the air in the borehole. Out of discussions about how pre-splitting works, we decided that perhaps there was some other explosive geometry which would make more efficient use of the explosive, and result in optimizing the pre-splitting process. Starting in the laboratory with small plexiglas blocks, we found that a fully-coupled explosive charge of 5 grains/in. of PETN gave random fractures extending about I in. from the borehole. A 1/4 in. plexiglas plug fit into a 1/4 in. hole in a similar block with the same fully coupled charge gave random fractures extending out only about 1/2 in. due to the energy absorbed in fracturing the plug. If the plug was split longitudinally but otherwise using the same geometry, it was possible to drive a fracture about 2 in. in either direction from the borehole, with an even smaller amount of fracturing around the borehole.

Jan 1, 1967

By B. W. Adams, R. W. Shilling

9.4-1. Basic principles. DESCRIPTION AND BASIC FUNCTIONS. The concept of inclined skip haulage is not new. The basic idea has been in use at least 400 years. The system is essentially a steeply inclined railway utilizing a hoist to pull the skips to the top or intermediate positions. Inclined skip haulage is often justified where high lifts would require long hauls for other haulage systems. The basic system types are as follows: 1. Single skip, single drum. This system is a single car pulled by a simple hoist system. 2. Balanced skip, single drum. A system utilizing two cars (skips on cables)--one overwound, one underwound-on a single drum. The dead weight of one skip and cable balances the dead weight of the other. As one skip ascends the other descends between two fixed points. 3. Balanced skip, double drum. This system is an elaboration of system No. 2, except the drums may be disengaged from one another to permit the skips to service various loading and dumping points. The preceding are major classifications subject to numerous variations. These variations may be in the skipway, skip design, loading methods, or unloading methods. A few examples follow: 1. The skipway may be standard or special railroad, an overhead monorail or cable, or a ramp for rubber tired vehicles. 2. The skip, per se, may be of varying dumping arrangements (side, bottom, etc.) or could be the box of the vehicle hauling to the skip. 3. Loading may be accomplished directly from the service unit to skip, through an intermediate loading device (e.g., a pocket), or into bins which feed the skip. Loading may be at more than one point. 4. Unloading of the skip can be simple or complex as required. Dumping may be to the haulage unit, to bins and feeders, or to a crusher and conveyor system. Dumping may take place at more than one point.

Jan 1, 1968

By Basil B. Zavoico

The production of all Russian fields incressed from approsimatctly 74,000,000 bbl. during 1926-27, to approximately 83,000,000 bbl. during 1927-28. Of this amount Baku was responsible for 54,.500,000 bbl., Grozny for 25,800,000 bbl., and Emba for 1,200,000 bbl. Accordingly, the proportional importance of the Grozny oil field has been considerably increased, the present production being twice that of 1913. The most notable factor of production during 1927-28 was the importance of gushers; 40.8 per cent. of the total was thus obtained. In the Baku fields 26 per cent. of the total, and in Grozny 74.2 percent'. of the total were produced by naturally flowing wells. This may be explained by the reduction in authorizations for drilling by the central authorities, and, accordingly, the local trusts increased the production to the planned figures at the expense of the proved prolific reserves. In Baku region, the Surnkany oil field produced 36.1 per cent. of the total Baku production, or an equivalent of 53,800 bbl. per clay, the largest percentage coining from the fifth horizon. Nest year's plans call for continual development of that horizon as well as of the sixteenth horizon in the Bibi-Eibat field, which was opened lately. In the Grozny fields the New Area (Bellik) WAS responsible for 79.4 per cent. of the total production. The authorized plan for 1928-29 production calls for 92,200,000 bid., and a reasonably accurate estimate for 1932--33 production is placed at. 156,000,000 bbl., with Baku contributing 101,800,000 bbl., Grozny 44,100,000 bbl., Emba 6,200,000 bbl., Kuban 2,100,000 bbl., Saghalien Island 1,400,000 bbl., and Russian Turkestan 400,000 bbl. In connection with recent discoverices in the Baku area these figures art. quite conservative. The average production per well per day in the Baku oil fields was 49.2 bbl., in Grozny 175 bbl.; in detail, Surakany oil field averaged 250 bbl., and the New Area in Grozny 514 bbl. per day. New completions in all Baku fields averaged 264 bbl. per day, and in Grozny 970 bbl. pel. day. These figures show a considerable progress in the technical developnlent of all fields, compared even with the prerevolution figures, since all the mprovements are noted in the fields that arc over 30 years old.

Jan 1, 1929

By Herty, C. H.

Three distinct processes take place during the conversion of iron ore to steel. First: the raw ore is reduced to metallic iron in the blast furnace and during this reduction certain constituents are dissolved in the iron, either through the solution of a solid into a liquid phase, as in the case of carbon, or by reduction of an oxide existing in the slag to the metallic element of that oxide, with subsequent solution of the metallic element in the iron. Thus by the process of solution and reduction, a pig iron results which contains about 4 per cent carbon, 1 per cent silicon, 1 to 2 per cent manganese, 0.1 to 0.4 per cent phosphorus, and 0.03 to 0.10 per cent sulfur. To obtain a sufficient yield of iron from the ore it is necessary that the reducing process be carried as far as possible, with the result that the impurities mentioned accompany the reduced iron. Second: to convert this iron into steel it is necessary to oxidize a part of the iron in eliminating the impurities to the extent desired for the particular product being made. In eliminating these impurities by oxidation a certain amount of iron oxide (FeO) dissolves in the steel. This dissolved FeO reacts with the metalloids, except sulfur, and to eliminate them to the desired extent it is necessary to leave a considerable amount of FeO in the metal. The reaction of iron oxide with silicon, manganese, and phosphorus results in the formation of the oxides of these elements. The oxides enter the slag and the extent of elimination of the elements depends on certain equilibria between the slag and the metal. The reaction of FeO with carbon results in the formation of the gases CO and CO, which, after saturating the metal, pass into the furnace gases. Third: after the iron has been purified to the desired extent as far as the elements listed are concerned, it is necessary to deoxidize the bath to remove the FeO, which is the vital factor in gas formation as the steel solidifies.

Jan 1, 1957

By Zay Jeffries

HOSTILITIES in Europe, Asia, and northern Africa were responsible for dislocations in rare-metal supplies during 1940. Although the consumption of some of the rare metals is small the dislocations may be relatively large. In some cases these are advantageous to North America and in others they are disadvantageous. Specific instances will be cited later. Barium T HIN films of barium metal on special steel balls greatly improve the life and reduce the friction of rotating anodes in highly evacuated small X-ray tubes and make possible the operation of these anodes at relatively high temperatures. These remarkable results are described in an article by Atlee, Wilson, and Filmer in the Journal of Applied Physics for September, 1940. Oil could not be used because of the high temperatures on the one hand and its volatility on the other. Fig. 3 shows diagramatically the rotating anode and the method of coating the steel balls with barium. The letters K.I.C. refer to "Kemet Iron Clad." The K.I.C. wire was mounted so that the barium could be evaporated by direct passage of the electric current. At high temperatures the barium dif-

Jan 1, 1941

By Gregory L. Gullord

In September, 1975, production was begun at the new milling facility of ASARCO Incorporated, the Young Mill. The mill is located adjacent to the Young Mine 35 km (22 miles) east of Knoxville, Tennessee. The Young Mill replaced the obsolete Mascot Mill which had been operating since 1913. The basic flow sheet of Mascot was retained for the construction of the Young Mill, with a 20% increase in capacity included in the design. The Young Mill is capable of processing 7710 - t/d (8500 stpd) of crude ore from three underground mines (Young, Immel and Coy) yielding 300 - t/d (330 stpd) of 63% zinc concentrate. The Young Mill is unique in that 92.5% of the ore milled is recovered as saleable products, including zinc concentrate, construction aggregate, sand and agricultural limestone. Processing at the Young Mill is divided into five basic stages: two-stage crushing, heavy media separation, grinding-flotation, concentrate filtering and drying, and limestone by-products recovery. The ore bodies are located in the Knox formation, which consists of primary dolomite, secondary dolomite and limestone with silicas in the form of chert and jasperoid. The zinc mineralization occurs as an exceptionally pure sphalerite. The ore contains an occasional abundance of pyrite and traces of fluorite and barite. TWO STAGE CRUSHING Ore enters the crushing section from two points. The Young Mine ore is hoisted into a 816 ton (900 st) bin which feeds a 1.52m (60 in) conveyor via a 1.52 m (60 in) vibrating feeder. The Immel ore is trucked 10 km (6 miles) by an independent contractor and weighed on automatic recording truck scales at the entrance to the property.

Jan 1, 1977

Industrial Preparedness To Cooperate with the U S War Department General Committee ARTHUR S DWIGHT, Chairman H FOSTER BAIN C K LEITH J EDWARD SPURR H. S MULLIKEN Alternate GEORGE OTIS SMITH POPE YEATMAN

Jan 1, 1923

By E. A. Almond

An attempt is made to explain the effect of stress on strain aging by examining the mechanism of yielding for a group of aged dislocations. The experimental results on which the theory is based indicate that a linear relationship develops between the aging stress and the discontinuous yield effect in a low carbon steel THE discontinuous yield effect that occurs in bcc metals after strain aging is usually explained by the interaction of interstitial atoms with individual dislocations. Attempts have been made to interpret the kinetics of strain aging in terms of interstitial segregation to nonrandom groups of dislocations1-3 but apart from Li's4 work little or no effort has been made to examine the effect of groups of aged dislocations on mechanical properties. It appears likely that such groups can be stabilized if a positive load is maintained on the specimen during aging5 and, furthermore, that the enhanced strain aging effect associated with aging under load might be due to the stability of these aged groups. The effects associated with this latter phenomenon have been described by Almond and Hull, Ref. 5, Figs. 2 and 3, and it is found that the upper yield stress, the lower yield stress, and the yield point elongation are increased by aging under load. The yield point elongation reaches a maximum value but the enhanced effect persists in the upper and lower yield stress values even after extended aging treatments when the general level of the flow stress curve rises. The flow stress, as measured at 8.5 pct total strain, however, is independent of aging stress. Almond and Hull5 showed that it was unlikely that the differences in mechanical properties could be caused by stress enhanced diffusion and they suggested that the effect was in some way associated with the different dislocation distributions that are obtained when specimens are aged with and without an applied stress. At that time no explanation was offered for the strengthening effect produced by stabilized dislocation distributions but additional tests have been performed to establish a quantitative relationship between aging stress and mechanical properties, and also to examine more closely the effect of varying the procedure for applying the aging stress. EXPERIMENTAL The material used was an iron wire containing 0.015 wt pct C, 0.002 wt pct N, and 0.006 wt pct 0. Tensile specimens with a 1 cm gage length and 0.08 cm diam were annealed at 850°C for 1 hr in vacuum to establish a grain diameter of 0.032 mm and then aged at 200°C for 24 hr. After this treatment the amount of carbon left in solution would be less than 10-4 wt pct, and ni- as aging time is increased. It is suggested that this observation, and effects that arise from varying the method of applying the aging stress, can be explained by a strengthening mechanism whereby dislocations are more difficult to move when they are aged in piled-up groups. trogen would be the main cause of strain aging. Tensile tests were performed in a hard beam machine at a constant crosshead speed of 0.02 cm per min and the specimen chamber was immersed in a temperature controlled silicone oil bath at 32" * 0.05"C. RESULTS All specimens were prestrained 5 pct before aging under stress and the results in Figs. 1 to 5 show the effect of aging time and aging stress on the following parameters ?UY = auy — ?F(5); i.e., the difference between the upper yield stress after aging,?uy, and the flow stress after prestraining 5 pct, ?f(5). ?LY = sly —sf(5); the difference between the lower yield stress after aging, ojy, and the flow stress after prestraining 5 pct. s8.5 = the flow stress at 8.5 pct total strain after aging at 5 pct strain. Varying the Loading Procedure. Three variations in the procedure for applying the aging stress were examined; i) After prestraining, the specimen was unloaded to a stress of 18 kg mm-2, aged at that stress, and then tested. ii) After prestraining, the specimen was unloaded to 2 kg mm-" then reloaded to 18 kg mm-', aged at that stress, and tested. iii) After prestraining, the specimen was unloaded to 18 kg mm-', aged at that stress, then unloaded to 2 kg mm- before testing. Specimens were unloaded or reloaded by decoupling a clutch in the drive transmission of the tensile machine. This enabled the crosshead to be driven manu-

Jan 1, 1970

By A. E. Bellis

There is a wide difference of opinion among rifle experts in the matter of barrel steel, and the relative importance to the life of the barrel of the steel's composition, heat treatment, structure, and physical properties. In order to promote discussion, which it is hoped may' lead to the discovery of the source of some of the traditional beliefs on the subject, the results of some tests are presented here, with data on heat treating, physical tests, and erosion. The detailed metallographic study of this steel is given in another paper.l The conclusions from these tests should be capable of wider application than their direct bearing on the barrel-steel problem. The Government specifications for barrel steel for model 1903 rifle (Springfield) call for carbon 0.50 to 0.60 per cent., manganese 1.00 to 1.29 per cent., silicon under 0.25 per cent., sulfur under 0.06 per cent., phosphorus under 0.08 per cent. The minimum physical requirements are: Tensile strength, 110,000 lb. per sq. in. (7733 kg. per sq. cm.); yield point, 75,000 lb. per sq. in (5472 kg. per sq. cm.); elongation, 20 per cent.; reduction of area, 45 per cent. Material fulfilling these requirements gives practically no failures when proof-fired, with charge of 70,000 lb. per sq. in. pressure, which is 40 per cent. in excess of the service charge. The turning and drilling operations show up any seams or laps such as accidentally occur even in the best heats of open-hearth steel, so that faulty stock is practically all eliminated before proof-firing. These physical requirements are met either by rolling billets at a proper temperature down to the form of tapered barrel blanks or by heat treating bars that have been upset to cnlarge the breech section. The transformation temperature of this steel, on heating, is around 1350" F. (732' C.), varying somewhat with the composition and the rate of heating. The actual treatment of the barrel blanks, as carried out by different manufacturers, gives four kinds of barrels, which will be called A, B, C, and D, for convenience. A barrels result from heat treating upset blanks by quenching at

Jan 1, 1920

By M. T. Abasov

The work suggests an approach to the problem of optimally planning the drilling and development of gas fields. The approach is based on systems analysis and allows a joint optimal choice of the total number of wells in the field, their spacing in the reservoir, the diameter of production tubing, drilling dynamics, the number and spacing of drilling platforms and well pattern schemes for each plat- form as well. The optimal solution satisfies the set of technical- economical requirements and provides a minimum value of discounted total expenditures (capital costs plus operating costs). A mathematical model of the problem is presented. This work represents an extension of the authors&apos; previous investigations that formulated and solved problems of optimal drilling of oil and gas deposits by directional wells [1] and problems of optimal control of the number of active wells in multi-reservoir oil and gas fields [2], [3]. For construction of the mathematical model we choose a range (Nmin, Nmax) of possible values of the total number of wells, N, which should be drilled in the deposit. Here Nmin, Nmax are minimum and maximum numbers of wells correspondingly, that should be drilled into the reservoir. N[€](Nmin = N1, ... N[o], ..., Nm = Nmax). For simplicity a one-reservoir gas field is considered. It is assumed that: (a) wells are spaced in the reservoir uniformly, (b) there is a certain algorithm providing a uniform spacing of the given number N[o] ([0]=l, . . . , m) of wells in the reservoir, (c) the diameters of the production tubing for all the wells are the same. The relation Qz (x), i.e. the dependence of the cumulative output of gas from the cumulative "well-years" of operations is constructed for each diameter dZ from the given range of diameters d1< ...< dr.

Jan 1, 1977

By H. Eyring, J. W. Fredrickson

Many theories have been advanced to explain the phenomena of elastic and plastic deformation. The object of this report is to present a mechanism for deformation, not radically differing from existing theories, but embodying.certain principles and concepts of these theories. This proposed theory cannot be the complete answer to the age-old problem of deformation but it is believed that it will be a step further in the proper direction. The plastic properties of solids have recently been reviewed by Seitz and Read.&apos; ZenerZ has reviewed and correlated the existing work on the anelasticity of metals. With these and others3&apos;4&apos;5 resent theories of deformation are well-known. Of the many theories now extant, those receiving most attention have been proffered by Griffith,6 Becker,l Smekal,8 Polanyilg Orowanlo and Taylor.&apos;&apos; Although differing quite widely in details, these theories are based more or less on a common principle of flaws or dislocations that exist in the crystal lattice. The effect of stress concentrations surrounding a flaw or crack has been treated most recently by Fisher and Hollomon.12 The various authors differ as to the exact nature of the flaws or dislocations, the causes of their appearance, or the manner in which they initiate slip. All of the theories agree that the number of dislocations increases with progressing deformation. It is universally agreed that the shearing process is not a simple gliding of ideally perfect atomic lattice planes. It is also agreed that the "macroscopic" process of gliding, or slip, is brought about by the spreading of a large number of locally initiated "glide steps" through the whole crystal. The occurrence of these "glide steps" is closely related to the presence of some kind of deviations from the "ideal" lattice structure, even in the undeformed crystal. These deviations are overcome &apos;&apos;one at a time," in a manner of speaking, due to some stress-concentrating effect.13 Proposed Theory Many solids show a continuously increasing deformation if subjected to a constant load of suitable magnitude in an extended range of temperature for a sufficient time. This deformation, or flow, may be said to be due, in general, to the tendency of the atoms or molecules in an elastically stressed body to readjust themselves in such a way that if the deformation were kept constant, a release of stress would take place. If rupture does not occur, this process of relaxation causes the rate of flow to obtain a final constant value, as shown in the curve (Fig I) obtained by Taylor and Quinney.14 In this state the release of stress by relaxation exactly counterbalances the increase due to the elastic strain on the test pieces.

Jan 1, 1949

By A. J. May

THE Star shaft is near the north boundary of the group of mining claims belonging to the Ground Hog Unit of the American Smelting and Refining Co., near Vanadium, N. Mex. The shaft bins and surface plant are served by a siding from the Atchison, Topeka and Santa Fe railroad, and this factor of railroad transportation influenced the choice of site. This paper describes the shaft sinking only, and does not cover the surface plant or other parts of the mine or its equipment. A. J. MAY is Engineer in Charge of Shaft Sinking, Ground Hog Unit, American Smeling and Refining Co, Vanadium, New Mexico. El Paso Meeting, October 1948. TP 2737 A. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Nov. 29, 1948; revision received June 22, 1949. Experience with this shaft seems of interest because of the mechanical mucking, good rate of progress, and the costs obtained at the present level of labor and material costs. The shaft has four compartments, as shown in fig. 1. Equipment: Riddell Mucker: The mucking was done by a clamshell bucket and a Riddell mucker, which, as in other shafts sunk with this machine, consisted of a carriage mounted on axles with double flanged wheels that traveled back and forth on a swinging track, parallel to the long axis of the shaft (fig. 2 to 6). Three air motors were mounted on the carriage; one hoisted the clam, one operated the bucket opening and closing device, and the third moved the carriage along the track. The carriage was designed with guide shoes and could be hoisted to the surface in either skip compartment without trouble (fig. 4). During blasting operations the clam was tied on the rails and the carriage was hoisted up a few sets to protect it from flying rock. No attempt was made to hoist the carriage and clam together lest the clam catch in the timbers. The track on which the carriage operated was built of extra heavy 6 in. pipe with 1x2 in. strap welded on top for rails. A Yzx4 in. angle was welded on the bottom of the 6 in. pipe with a 1 in. round rod welded in the angle for a filler and additional strength. The end or cross pieces of the track which kept the track to gauge, were also made of reinforced, heavy 6 in. pipe and were bolted to the long members on which the carriage traveled (fig. 2 and 3). This section of track was used, instead of heavy railroad rails, for two reasons: (1) this section would bend before it snapped and was much safer for the men in the bottom; and (2) it was a better blasting shield, and protected the timbers better, in fact, only five wall plates were broken by blasting in 1926 ft of shaft, although 150 to 225 lb of 40 pct powder was loaded per round, depending on the ground conditions. The swinging track was supported by four % in. hoisting cables anchored to the end plates. There were also four safety cables fastened two to three sets above the carriage, which would catch the carriage if one of the supporting cables should fail for any reason (fig. 7). Clamshell Buckets: Two clams were used, each of % yard capacity, both manufactured by Blaw-Knox, Nos. 662 and 663. The No. 663 clam was 4 in. wider and 5 in. shorter than the No. 662, but seemed to dig and load more satisfactorily. It was easier to hoist and lower when it came up for repair. It did not turn over as easily, had less spill

Jan 1, 1951

By A. J. May

THE Star shaft is near the north boundary of the group of mining claims belonging to the Ground Hog Unit of the American Smelting and Refining Co., near Vanadium, N. Mex. The shaft bins and surface plant are served by a siding from the Atchison, Topeka and Santa Fe railroad, and this factor of railroad transportation influenced the choice of site. This paper describes the shaft sinking only, and does not cover the surface plant or other parts of the mine or its equipment. A. J. MAY is Engineer in Charge of Shaft Sinking, Ground Hog Unit, American Smeling and Refining Co, Vanadium, New Mexico. El Paso Meeting, October 1948. TP 2737 A. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Nov. 29, 1948; revision received June 22, 1949. Experience with this shaft seems of interest because of the mechanical mucking, good rate of progress, and the costs obtained at the present level of labor and material costs. The shaft has four compartments, as shown in fig. 1. Equipment: Riddell Mucker: The mucking was done by a clamshell bucket and a Riddell mucker, which, as in other shafts sunk with this machine, consisted of a carriage mounted on axles with double flanged wheels that traveled back and forth on a swinging track, parallel to the long axis of the shaft (fig. 2 to 6). Three air motors were mounted on the carriage; one hoisted the clam, one operated the bucket opening and closing device, and the third moved the carriage along the track. The carriage was designed with guide shoes and could be hoisted to the surface in either skip compartment without trouble (fig. 4). During blasting operations the clam was tied on the rails and the carriage was hoisted up a few sets to protect it from flying rock. No attempt was made to hoist the carriage and clam together lest the clam catch in the timbers. The track on which the carriage operated was built of extra heavy 6 in. pipe with 1x2 in. strap welded on top for rails. A Yzx4 in. angle was welded on the bottom of the 6 in. pipe with a 1 in. round rod welded in the angle for a filler and additional strength. The end or cross pieces of the track which kept the track to gauge, were also made of reinforced, heavy 6 in. pipe and were bolted to the long members on which the carriage traveled (fig. 2 and 3). This section of track was used, instead of heavy railroad rails, for two reasons: (1) this section would bend before it snapped and was much safer for the men in the bottom; and (2) it was a better blasting shield, and protected the timbers better, in fact, only five wall plates were broken by blasting in 1926 ft of shaft, although 150 to 225 lb of 40 pct powder was loaded per round, depending on the ground conditions. The swinging track was supported by four % in. hoisting cables anchored to the end plates. There were also four safety cables fastened two to three sets above the carriage, which would catch the carriage if one of the supporting cables should fail for any reason (fig. 7). Clamshell Buckets: Two clams were used, each of % yard capacity, both manufactured by Blaw-Knox, Nos. 662 and 663. The No. 663 clam was 4 in. wider and 5 in. shorter than the No. 662, but seemed to dig and load more satisfactorily. It was easier to hoist and lower when it came up for repair. It did not turn over as easily, had less spill

Jan 1, 1951

By William T. Folwell

The Lucky Friday mine east of Mullan, Idaho, is an outstanding example of a property in the Coeur dlAlene district where a small and insignificant-appearing silver-lead-zinc vein at the surface has changed at depth into a large vein of great importance. The Lucky Friday vein has little if any surface expression and above the 1200 level the ore shoots are small and discontinuous. Between the 1200 level and 2450 level, the lowest developed level, the main ore shoot has shown remarkable improvement on each succeeding lower level, and today the mine is one of the major lead-silver producers in the Coeur d&apos;Alene district. History: The Lucky Friday property is on the north side of the South Fork of the Coeur d&apos;Alene River in sections 25, 26 and 35, T. 48 N., R. 5 E., Hunter mining district, Shoshone County, Idaho. This is about one mile east of the town of Mullan, which serves the eastern portion of the Coeur d&apos;Alene district. The southern part of the property is crossed by a branch line of the Northern Pacific Ry. and by U. S. Highway 10. This highway is the principal road crossing the panhandle of Idaho and connects the district with Spokane, Wash., on the west and Missoula, Mont., on the east. The main portal and surface plant of the Lucky Friday mine is at an elevation of 3365 ft, only a short distance above the valley floor and a few hundred feet from U. S. Highway 10. so the mine is readily accessible for year-round operation. The property is comprised of six claims, known as the Lucky Friday group, owned outright by the Lucky Friday Silver-Lead Mines Co. There are four patented claims, Good Friday, Lucky Friday, Northern Light, and Lucky Friday Fraction No. 2 (Mineral Survey No. 3028), and two unpatented claims, Hunter and Creek. In addition, Lucky Friday owns an undivided one-half interest in the Hunter Creek property. which adjoins the Lucky Friday group on the north: a 90 pct interest in the mineral rights in the Jutila Ranch (160A), which adjoins the Lucky Friday group on the east; and a 60 pct interest in the Lucky Friday Extension claim group, which adjoins the Lucky Friday group on the west. The company also has a long-term mining lease on the Hunter Ranch, which adjoins the Lucky Friday group on the west. The claims of the Lucky Friday group were located between 1899 and 1906. The Lucky Friday Mines Co. was organized in 1906 and did considerable exploration work by surface trenching and shallow underground workings, only to see the property sold by the Shoshone county sheriff to satisfy labor claims totaling $2000 in 1912. Another firm, Lucky Friday Mining Co., bought the claims in 1914 and spent 12 years driving what is now known as the tunnel level crosscut. This tunnel intersected a vein previously exposed in a higher tunnel, but it was only a few inches wide. The vein was followed a short distance westerly but was so unpromising that the work was discontinued in favor of extending the main crosscut tunnel several hundred feet north. No ore was found and all work was discontinued. The property was held in such low esteem by the firm that it let taxes amounting to less than $15 a year go delinquent for nine years. Then the property lay idle for two more years until in 1938 John Sekulic, a Mullan service station operator, took a lease, with a $15,000 purchase option, on the advice of an old miner who had worked in the mine. Sekulic re-opened the tunnel level crosscut and explored the vein with an easterly drift for about 200 ft. The vein was too narrow to be of commercial value but was believed interesting enough to warrant further exploration at depth. Lacking funds to explore the vein at depth, Sekulic tried to get the district&apos;s larger operating companies to take over his lease and option. They were not interested because of the lean tunnel level showing and the fact that the property lay between the White Ledge fault on the north and Osburn fault on the south, an area which geologists always considered unworthy of exploration. Sekulic then organized the present company and assigned his lease and option to it for stock. This was in 1939. Enough stock was sold locally to finance sinking of a shaft 100 ft from the tunnel level east drift. The vein at this additional depth still was not commercial but showed some improvement. Treasury stock was offered at 5 to 10C a share

Jan 1, 1959