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By C. A. Stewart

I. INTRODUCTION AND GENERAL STATEMENT OF RESULTS. The field-work upon which this paper is based was done in the summer of 1910, and was made possible by the courtesy of the Imperial Copper Co., which gave me access to its property. For their kindness in this respect I wish to express my thanks to W. F. Staunton and Meade Goodloe. Moreover,. B. F. Smith and G. B. Gentry, mining engineers in the employ of the company, gave invaluable assistance by placing maps,. etc., at my disposal, and by discussing with me the problems of the district; and it is a great pleasure to acknowledge this indebtedness to them. In order that the significance of some of the details that follow may be better appreciated at first reading, a general outline of the conclusions reached is given here. In the Silverbell mining-district a series of Palaeozoic limestone blocks is completely surrounded by post-Palaeozoic igneous rocks, intruded in the following order: alaskite, alaskite-porphyry, biotite-granite, andesite, and quartz-porphyry. The biotite-granite is believed to represent a differentiation of the. magma from which the alaskite-porphyry came ; and this conclusion-has an important hearing on the origin of the ores. The intrusion of both alaskite-porphyry and biotite-granite was followed by the emission of magmatic waters, which sericitized and silicified the alaskite-porphyry and granite, and produced in the limestone, by the addition to it of silica, iron, and alumina, great masses of garnet, quartz, and wollastonite. Following close upon these solutions came metal-bearing magmatic waters, which impregnated porphyry, granite, and alas kite with cupriferous pyrite, and deposited in the garnet zones chalcopyrite and copper-bearing pyrite that make important bodies of contact-metamorphic ores. Secondary enrichment has taken place in the disseminated ores in the igneous rocks,. raising their copper-content to 2 per cent. in some cases,. and extensive drilling has been undertaken to block out these deposits. East of the contact-metamorphic deposits there is a fissure-vein of lead-silver ore in the quartz-porphyry. This rock is distinctly younger than the alaskite-porphyry and the granite; and the lead-silver ores belong, therefore, to a period of mineralization distinctly later than the copper-deposition.

May 1, 1912

By H. O. Hofman

WHEN dry chlorine gas is made to act in the cold upon finely¬divided gold,' it converts the latter with evolution of heat into auro-auric chloride, Au2CI4, a hard, dark-red, hygroscopic salt. Moisture splits this salt into aurous and aerie chloride, Au2C14 = AuCI + AuCI3; treatment with water converts it into auric chloride and gold, 3 Au2CI4 = 4 AuC13 + Au,. Aurous chloride, when stirred with water, undergoes a similar decomposition, 3 AuCI = AuCI3 + Au,. These decompositions of auro-auric chloride and of aurous chloride furnish the explanation for the practice of moistening an ore before it is treated by the Plattner chlorination-process; practical experience having shown that gaseous chlorine gave an unsatisfactory extraction with dry ore. According to Rose,' fine gold is acted upon more slowly by chlorine than gold containing some base metal, e.g., copper. He also states that small quantities of silver increase the rate of solution, but adds that the coating of silver chloride formed checks, and finally stops, further action, if the percentage of silver be increased beyond a certain undetermined amount. Bromine' acts upon finely-divided gold in a manner similar to chlorine; the auro-auric bromide, Au2Br4, however, is not hygroscopic. Upon treatment with water it is decomposed as is the corresponding chlorine salt. It is generally believed that bromine does not act as energetically upon gold as does chlorine. This seems to be borne out by the results in the leach-

Mar 1, 1905

By Richard W. Smith

Why are we losing our liberties? (1) . . . because our local chambers of commerce come to the National Chamber's annual meeting, vote for a policy on federal economy, and then go to Capitol Hill and ask for a federal project for their area. (2) ... because our mining men ask for a subsidy to mine submarginal deposits that can't be economically mined in a free economy. (3) . . . because 18,000 men and organizations all over the country recently wired or wrote a Congressional committee asking that the Army Engineers be exempt from the President's power to reorganize the Executive Branch of the government under the recommendations of the Hoover Commission.

Jan 1, 1949

By R. F. Shurtz

R. Duval (Mining Engineer, Ancien eleve de PEcole Polytechnique, Paris, France) I do not agree with the Eq. 3, reading: m =1/100- [(0.214x30.4) + (0.7B6 x0.00)] =6.5pct CaO If 0.214 and0.786 were proportions by weight, the equation would represent the well known mixtures law of the conventional arithmetics and 6. 5 pct CaO would be the correct average content. But it is not the case as the author states: "In samples consisting of single grains of mineral, those grains must, as already mentioned, be either of dolomite or magnesite. Since 78.6 pct of the deposit consists of magnesite and 21.4 pct of dolomite (excluding for present purpose the presence of other minerals), for any single grains picked at random the probability will be 0.214 that is it dolomite and0.786 that is it magnesite. In 1000 such samples the expected numbers of dolomite and mapesite grains will be 214 and 786 respectively." 0.214 and 0.786 would be proportions by weigbt under the necessary condition that all grains of dole mite and magnesite should have an identical weight. Obviously it is not the case, as the specific gravities are not the same for mapesite and dolomite and the volumes of the grains are different. Furthermore, because of these differences the conditions for a random sampling are not fulfilled and we are not authorized to state that the probabilities are, respectively, 0.214 and 0.786. The author however makes a simple application of Eq. 1: M = 1/n— ? fi x i . n Should we deduce that this relation is wrohg? Not at all, but when applying Eq. 1 you must not overlook what it actually. means. Eq. 1 gives a definition of the arithmetic mean of a total of n observed values Xi and nothing else. But the average conteht of a deposit has not the same significance. It is the ratio between the weight of concerned mineral in the deposit and the total weight of the deposit. As from 1000 particles the 214 of dolomite and the 786 of magnesite have not the same weight, the two definitions do not concur, and when applying Eq. 1 the result is an arithmetic mean of figures which has no connection with what is named average contentof a deposit. The situation is similar to the calculation of an average velocity. If a car travels a first mile over at 30 miles per hr and a second mile over at 60 miles per hr, when applying formula 1 you find as average velocity for the 2 miles: 30+60 ------- - 45 miles per hour. Many people calculate in this way and they do not realize that a mistake is involved. In fact the definition of he average velocity for the 2 miles is the quotient of the distance of 2 miles by the time (in hours) necessary for 2 miles travel, i.e.: 2 ---------- = 40 miles per hr. 1 + 1 30 60 In other words, the average volocity wanted is not the arithmetic but the harmonic average of the two velocities. The above mentioned bias in the calculation of the average contents of deposits is frequent, even in the assessments made by experienced engineers and is independant of what is named the sampling error. In order to supress the bias and to be able to use Eq. 1, you must apply a correction. An example on the subject can be found in an article by Duval et al. in the January 1955 issue of the ''Annales des Mines" (French), page 19. R. F. Schurtz (Author's Reply) Mr. Duval's position is quite correct. The proportions shown for dolomite and magnesite., respectively, of 0.214 and 0.786 are, in fact, proportions by weight uncorrected for specific gravity. In our day to day operation of producing magnesite from these mines at a very substantial rate, we do not normally make corrections for the difference between the specific gravity of dolomite and that of magnesite. If these corrections are made in Eq. 3 as shown in my article, then the numbers of grains turn out to be in proportions of 0.226 dolomite and 0.774 mapesite instead of the values actually shown in the equation. For the purposes of our work, and in view of the inherently low accuracy of the data, this correction was not deemed worthwhile making.

Jan 1, 1961

By M. Yokoi, K. Yoshida, T. Ishihara

The solubility of aluminum in cryolite melts has been determined by measuring the weight loss of aluminum submerged in melts contained in closed silicon nitride crucibles. The solubility in pure cryolite has also been checked with radioactive isotopes. The solubility decreases with increasing A1203 concentration, decreasing temperature, and decreasing cryolite ratio. The change of aluminum solubility in pure cryolite with the activity controlled by using Al-Ag and A2-Cu alloys supported the assumption that aluminum would dissolve in the form of monovalent ions according to the reaction 2AI + A1' —3 A1'. DURING the electrolytic production of aluminum, the electrodeposited metal redissolves in the bath in some form and is subsequently oxidized back to the former ionic state. This dissolution is believed to cause a decrease in current efficiency. Reported data on the solubility of aluminum in cryolite melts are scattered. The mechanism of the dissolution has not yet been determined, although certain tentative interpretations have been proposed. Many previous studies were performed under different test conditions. Studies were performed in open Systems and closed systems,'o and some were apparently not given sufficient time to Saturate." The solubility can be determined either analytically or gravimetrically. The analytical determination involves measurement of hydrogen gas evolved when frozen cryolite bath saturated with aluminum is treated with sodium hydroxide or hydrochloric acid. The gravimetric determination, adopted by many investigators,'-37'3>14 involves measurement of weight loss of the metal after equilibration with a known amount of salt. It appears that the former technique yields much smaller values of solubility than the latter technique. This large difference in results strongly suggests that the dissolved metal in the cryolite melt is no longer in the metallic state. In the present study, the solubility was determined by the gravimetric technique at various activities of aluminum using A1-Ag and Al-Cu alloys. An attempt was also made to determine the ionic (or atomic) state of the dissolved metal from the solubility data. 1) EXPERIMENTAL Apparatus. Preliminary tests demonstrated that silicon nitride was suitable as a container material since it is compatible to both the fluoride melts and molten aluminum. Crucibles were made from silicon powder nitrided with Nz gas at 1600°C. The good compatibility between crucible and melts is due to the near perfect nitrogenation of silicon. Only small amounts of silicon were found in the aluminum metal after the experiments. The crucible dimensions were 90 mm inside height, 38 mm ID, 105 mm outside height, and 50 mm OD. Close-fitting lids made of silicon nitride were also prepared. The crucible was placed in a close-fitting graphite crucible heated by a vertical Kanthal furnace, Fig. l. The lid and crucible are fastened tightly together by the condensation of vaporized bath in the narrow gap between them. The temperature was measured in the wall of the graphite crucible and controlled within * 5°C. The temperature difference between the measuring point and the interior of the melts was initially determined. Materials. The materials used were: Greenland cryolite (impurities CaF,, 0.5 pct; AlzO3, 0.2 pct; high-purity aluminum metal (purity 99.99 pct); A1203 and AlF, (provided by Nippon Light Metal Co., Ltd); reagent-grade NaF and CaF2; and 99.9+ pct Ag and Cu. Powdered reagents were carefully dried before use. A1-Ag and A1-Cu alloys were prepared before the experiments; their composition was checked by analysis

Jan 1, 1969

By Akira Ishizu, Nobuo Matsui, Tatsuya Nagahama

In the concentrator of Kamioka Mine, lead-zinc bulk differential flotation is practiced. It results in three kinds of thickener over- flow, namely; tailing, lead concentrate and zinc concentrate, which are discharged as wastewater. Since these three kinds of overflow differ from each other in water quality, they are separately recycled to appropriate processes based on their water quality. The tailing and the lead concentrate thickener overflow are directly recycled to their respective suitable processes. However, as the zinc thickener overflow contains Cu(CN)3--, ZnS, CdS, CN-, xanthate and frother, the total zinc thickener overflow cannot be recycled at one time to any one process without creating harmful effects. Consequently, all the ingredients in the zinc thickener overflow other than CN- are floated by Ion-Precipitate-Ultrafine-particle flotation, using Nagahm flotation machine. After this treatment, the zinc thickener overflow, containing only CN-, is effectively recycled to the lead-zinc differential flotation circuit. Thus the integrated recycling of waste-water greatly reduces environmental problems, and in addition it generates a sizeable profit by improving overall mill recovery, and permitting the reduction of reagent consumption.

Jan 1, 1976

By JOHN R. SUMAN

TONIGHT I want to speak of the current problems and the postwar difficulties facing American industry. American industry has done an outstanding job in adjusting its operations to wartime necessity. The national income for 1942 was nearly $120,000,090,000 and of this outlay approximately $55,000,000,000 went for war material. The enormity of the expenditures for war material is indicated by the fact that the expenditures for war purposes alone were greater than the annual gross national income in the period 1931-34. This tremendous production job in turning out a record output of war products was accomplished by industry in spite of a large number of restraining and hampering influences, and I want to mention several of the more important problems.

Jan 1, 1943

By M. D. Cooper

For the purpose of obtaining some first-hand data in regard to the shooting down of coal in bituminous mines, it was the writer's good fortune to be employed as a shot-firer for almost one year. In all, 6020 shots were fired by him during this period. All of the work was done in what is known as the South Main section of the No. 2 Mine of the Ellsworth Collieries Co., Ellsworth, Pa., a subsidiary of the Lackawanna Steel Go. Position of Shot-firing in the Cycle of Operations The established practice of the Ellsworth Collieries Co. calls first for the undercutting of the coal by electric chain machines across the face of the entry or room. This is followed by replacing any posts removed by the machine-runners, or the setting of new posts or timber, as may be required. After this has been done the machine cuttings are loaded into cars. Then the coal is spragged, and a shot hole drilled to the depth of the undercut. Next in order, the shot-firer loads the hole and fires the charge. The working place is then cleaned up, timbered if necessary, the sides and face squared up, and loose material taken down. The face is again ready to be undercut. Materials Used Permissible explosives were used exclusively in blasting coal. These were Carbonite No. 2, a nitroglyccrine explosive; and three different makes of ammonium nitrate explosives, Tunnelite, Red H, and Mine-ite 5-D. The last-named was used during the greater part of the period. It is interesting to note that even at the beginning of this work, the gases resulting from the explosions of the charges caused no especial distress on. the shot-firer's part and this continued to be the case except on the occasions when a change was made from one class of explosive to another.

Jan 1, 1918

By Paul Rhedey

Various commercial pelroleum cokes were heat-1,reated at temperatures between 500° and 1500°C, in a nitrogen atmosphere, in laboratovy induction furnaces. The rate of tenlperature rise was varied betzveen 10" and 300°C per min, or the green cokes were flash-calcined, orv a combination of heating rates was used. Changes in the rate of heating had only negligible effects on the degree of' calcination as characterized by mean crystallite thickness and the chemical properties of the calcined coke. The physical structure of the cake was however significantly affected when rate of heating during calcination exceeded 50°C per min over the temperature range of 600° to 900°C. Surface-accessible porosily increased with the .rate of temperature rise, and this was accompanied by a change in pore size distribution. Source and properties of the green coke also had an influence on the structure of the calcined coke. The evidence presented suggests a similar mechanism of porosity development in petroleuiiz coke during calcination in industvial equiplnenl, such as rotary kilns. An increase in surface accessible povosity incveased the pitch binder requirement when the coke was used as aggregate in Soderberg paste. A correlation was established between calcined coke porosity and paste binder requiremenl. ManY results have been published on the changes of properties of petroleum coke during calcination, such as chemical composition, real density, electrical resistivity, crystallite and pore structure. The correlation of these properties with temperature of calcination and time at maximum temperature has been rather well established in both laboratory and pilot plant experiments. Surprisingly little attention has been given however to the effect of calcination conditions, such as rate of temperature rise or furnace atmosphere, on the chemical and structural properties of the calcined coke. It has been observed that petroleum coke, when calcined in industrial equipment, acquires higher porosity and lower real density than those attainable in laboratory furnaces at apparently identical calcina- tion temperature and soaking time. This paper describes a study of the effect of rate of heating during calcination on calcined coke properties using green petroleum cokes of different volatile matter, hydrogen, and sulfur content. An attempt was made to correlate the changes in coke structure with the flowability of anode paste of the type normally used in aluminum reduction cells. EXPERIMENTAL Petroleum Cokes Used. In the study of release of volatile matter and sulfur during calcination and the effect of rate of heating on calcined coke properties two delayed cokes of different sulfur contents were used. The results of analysis of the green cokes are given in Table I. In the study of the effect of flash calcination thirty-six commercial petroleum cokes from twelve different refineries were used with a range of properties shown in Table 11. Calcination Conditions. Calcination experiments were carried out in a laboratory induction furnace. In each run a 200-g sample of dry green coke sized to 10 by 65 Tyler mesh was calcined in a graphite crucible. The crucible containing the sample was placed in the middle of a stack of eight others filled with metallurgical coke to reduce temperature gradients within the sample. Temperature was measured by two Pt, Pt-10 pct Rh thermocouples and controlled by a Celectray instrument. The two couples generally agreed within 5°C. In the study of volatile matter and sulfur release the samples were heated to temperatures in the range of 500" to 1500°C at a rate of 10°C per min in a nitrogen atmosphere and held at the final temperatures for 30 min. In the study of the effect of heating rate on calcined coke properties the desired rates between 10' and 300°C per min were obtained by manually adjusting the power input. Flash calcination was carried out by dropping 100 g of the green coke into the graphite dish preheated to the calcination temperature. Because of the small heat capacity of the furnace the coke was introduced in 20-g portions at a time. For this purpose a 2-in.-long nipple between two 1-in. gate valves installed on the top flange of the furnace served to provide a gas seal while feeding coke to the furnace. It was estimated that the temperature of the coke reached that of the furnace at a rate of approximately 1000°C per min. Holding time at final temperature was also 30 min. Calcined Coke Proper- Determined. Porosity was determined on 20 by 35 Tyler mesh samples using an Aminco-Winslow mercury pressure porosimeter with an operating range of 1.8 to 3000 psi absolute pressure (100 to 0.05 p pore diameter range).' Apparent density was obtained by the mercury poro-simeter. It represents a particle density of the 20 by

Jan 1, 1968

By AIME AIME

MINERALS and mineral resources are recognized as one of the things that nations are prone to quarrel about. The territory that was arbitrarily incorporated into the Polish Republic after the World War produces a variety of essential minerals on a substantial scale. The accompanying table has been compiled to give some idea of the relative importance of this production. Poland, before the recent war started, had an area of 150,000 sq.mi., just a little larger than the State of Montana. Its population was 32,000,000, equal to that of New York, New Jersey, Pennsylvania, Connecticut, and Massachusetts. Though not a highly industrialized country it nevertheless has important mines, steel plants, petroleum refineries, and sundry metal-fabricating manufactories. Coal, petroleum, iron and steel, zinc, lead, pyrite, potash, and salt are produced in excess of the country's domestic consumption. It has no appreciable production of copper, chromium, tungsten, nickel, tin, mercury, manganese, aluminum, magnesite, fluorspar, or graphite.

Jan 1, 1939

By F. J. Radd, R. L. McGlasson

The mechanisms of corrosion and corrosion fatigue damages to sucker rods are examined from a fundamental electrochemical viewpoint, and the relationships of sucker rod microstructures to these damaging effects are defined. The common corrosive agents in crude oil production and storage are discussed as to types of attacks that they cause. The variations of fatigue strength with alloy content, surface preparation, and stress ranges are analyzed. Finally, conclusions are drawn as to how microstructures largely influence and control sucker rod performance. INTRODUCTION Greater than 90 per cent of the crude oil produced by artificial lift methods is produced through sucker rod pumping systems. In the USA each year more than $14 million is spent for sucker rods in new and old pumping installations. It is important to make this investment a wise one. Essentially, the sucker rod system is a long series of steel rods, operating under alternating stress loadings and under mechanical vibrations, which connect the surface power source with the subsurface pump. These rods often operate in a variety of corrosive media, including water, brines, carbon dioxide, hydrogen sul-fide, and air. Basically, this paper is concerned with the problem of the behavior of certain steels in various corrosive media under high and dynamic stress patterns. The main objective of this paper is to show how, once the stress effects and the corrosive environments are defined, one can still realize structural control and beneficial influence upon corrosion and corrosion fatigue. FUNDAMENTAL DEFINITIONS AND ANALYSIS In the corrosion and stress behavior of this solid state mixture of iron and iron carbide called steel, one must analyze upon what major factors does its service life depend and of what, exactly, is steel composed. The former question can be referred to two external factors and two internal factors. The two major external factors are: 1. the nature of the stress patterns — size and range of stresses present 2. the nature of the corrosion — erosion environment — the corrodent and/or wear involved The two fundamental internal factors that govern the strength and useful life of a steel both in normal duty and in heavy-duty corrosion service are: 1. the nature of the steel volume properties 2. the nature of the steel surface properties For the latter question normal steel is defined as a solid mixture of iron and iron carbide particles, generally with certain accidental and intentional elements present. (The amount of iron carbide present is determined by the carbon content of the steel.) These additional elements generally dissolve into either the iron (ferrite) or the iron carbide (cementite) components.

Jan 1, 1955

By R. G. Bentsen, D. A. T. Donohue

The cyclic steam injection process has become the most widely applied and most successful thermal recovery technique in use today.1-4 Normally, steam stimulation is repeated several times during the life of a project. The problem is to select the optimum number and size of steam treatments so that the profits realized over the entire life of the project are maximum. Obviously, operating costs rise directly with the frequency and amount of steam injection. Operating efficiency, on the other hand, can be adversely affected if the amount and frequency of stimulation are insufficient. Thus arises a multistage decision problem; that is, a decision must be made at each sub-period (day, for example), as to whether steam should be injected, and, if so, how much. It is desirable that these decisions be made in such a way as to secure maximum net profit before taxes. This can be achieved by dynamic programming. To use dynamic programming, we must be able to predict not only the production rate-time curve resulting from a given stimulation, but also how changes in stimulation policy will affect this curve. In this study, the effect of steam injection on well performance is estimated on the basis of a simple model that takes into account both steady-state and transient flow. Oil production rate is calculated for both types of flow, and the larger of the two values is used. This allows the calculation to change to radial steady-state flow at such time as this mechanism becomes more effective than the transient mechanism. We propose to show how dynamic programming can be used to optimize the steam soak process with respect to net profit. For this, a mathematical representation of the steam soak technique is necessary. A further objective, therefore, is the formulation of a mathematical model capable of realistically simulating the physical process. Oil Production Response Model The advantage of using a mathematical model is its generality and ease of manipulation. However, the model must be an accurate representation of the physical process, for a solution derived from a model can be no better than the model itself. The question now arises as to how well the model should be expected to simulate reality. In general, the simplest model should be tried first, for additional accuracy is apt to require additional cost and time. Also, the search for a high degree of accuracy may yield only greater complexity and difficulty without yielding significantly better results. With this in mind, we kept our model as simple as possible, yet generally representative of the actual physical process. The number of state variables is limited to three, and stochastic variations in the parameters are ignored. As a consequence, use of this model should be restricted to cases in which the underlying assumptions are not severely violated. In

Jan 1, 1970

By E. Northcott, F. N. Oberg

High-grade concentrates from Florida pebble phosphate deposits have long been recovered by wet methods. When work was begun in the field, only the coarse pebble (generally +1 mm) was recovered. In the 1920's a froth flotation process was developed to separate quartz from the finer phosphate fractions, but this method proved unsatisfactory for treating the intermediate fraction (-1 mm +35 mesh). Consequently various other processes were adapted, including agglomerate tabling and spiral separation of reagentized feed. Today the three processes are used-washing and screening for coarse pebble, tabling or spiraling for intermediate sizes, and flotation for -35 mesh material.

Jan 10, 1958

By J. S. Miller

THIS paper describes the method of mining a steeply pitching anthracite seam on a heavy pitch in the Orchard vein in No. 1 Tunnel of the Lehigh Coal and Navigation Co. The Orchard vein in the Nesquehoning district has a thickness of about 15 ft. on the level of the gangway, which is sufficient thickness for the gangway to be driven full width in coal. Above the gangway, the vein becomes thinner and near the surface is about 8 ft. thick. The coal is of good hard quality and carries a fairly good roof. The pitch of the vein is between 65° and 70°. Usually the breast-and-pillar method has been employed in the mining of the steeply pitching veins in this section. This method produced good results as far as the mining of the breast was concerned but the recovery of the pillar was difficult as to full extraction and expensive as regards costs. To recover these pillars after the breasts were driven, three methods were commonly employed; they were known as the pillar-breast method, pillar-skipping method, and pillar-chute method. The pillar-breast method consisted in putting a battery in one corner of the pillar (Fig. 1), coupling the manways A and B of the two adjacent breasts for two or three lengths ahead of the working face of the pillar being cut, and working the pillar as a breast. This method was used by the more skilled miners of the earlier days but has practically been abandoned. A slight variation of this method is used to some extent. In this case, the pillars are left wider than ordinarily and of sufficient width for the full-width breast to be driven in and a small pillar left on each side of the pillar breast between it and the regular breasts. With the completion of the "pillar breast," the ribs are drilled full of holes, loaded, and fired.

Jan 3, 1925

By M. H. Pilkuhn, H. Rupprecht

The junction luminescence of GaAsxP1-x diodes containing up to 47 pct Gap was studied. Diodes were prepared by diffusing zinc into n-type material which was eithw boat- or vapor-gvown. Observations concerning the energy and the line shape of the spontaneous emission are presented and discussed; in particular the influence of inhomogeneous phosphorus distribution is considered. The role of reabsorption of the junction radiation is pointed out. With increasing current, the peak emission shifts to higher enwgy. This is alm true in cases where slow line narrowing occurs simultaneously. The length dependence of the threshold current density for lasing at 77 °K was examined for lasers made from the same material. It was found that longer lasers have lower threshold curvent densities, Estimates of the characteristic laser losses are given. Data concwning the range of Gap concentration in which stimulated emission was observed are presented. FORWARD-BIASED p-n junctions in mixed alloys of GaAsxPl-, emit light whose wave length depenfs on the composition of the alloy ranging from 8400A (x = 1) to 5350A (x = 0) at 77°K. Laser action in vapor-grown material was first reported by Holonyak and evacqua,' and in boat-grown material by Ainslie, Pilkuhn, and Rupprecht.' In both cases this was observed for low Gap concentration where GaAs,Pl-, is still a "direct" semiconductor with the lowest conductance-band minimum at k = 0 and a subsidiary minimum in the [loo] direction. As the Gap concentration is increased, the subsidiary [loo] minimum in the conduction band is lowered and finally becomes the lowest minimum for Gap concentrations above 50 pt,3 making the alloy an ''indirect" semiconductor. It is of interest to determine up to which Gap concentration stimulated emission can be observed. EXPERIMENTAL Boat-grown material was prepared by N. Ainslie of IBM by horizontal Bridgman technique. The material was n type and silicon-doped up to a concentration of 10'' cm-'. Vapor-grown material was made by E. Hull of IBM or received from Merck & Co. Diodes were fabricated in a similar manner as in the case of GaAs by zinc diffusion. Details about the diffusion can be found in a later paper.4 All diodes were of Fabry-Perot structure with two parallel sides cleaved and the other two sides sawed. Their width ranged from 50 to 150 p and their length from 50 to 700 p. RESULTS AND DISCUSSION Spectral Distributions. A typical display of the junction luminescence of a GaAsxPl-, diode is shown in Fig. 1. The liquid-nitrogen and room-temperature curves which were taken at a very low current (25 ma dc) show a relatively sharp high-energy emission which is band-edge radiation. They also show a broad low-energy emission which is due to transitions to deep impurity levels. This latter emission line may have different shapes and positions depending on the nature of the material. It can sometimes be quite near the band gap giving rise to "bumps" in the high-energy line. The results presented in this paper will be concerned only with the high-energy emission which becomes more intense relative to the low-energy emission when the current is increased and the temperature is lowered.

Jan 1, 1964

By H. Conrad, V. V. Damiano, G. J. London

The slip mode activated during the c axis compression of single crystals of commercial-purity ingot SR beryllium, high-purity (twelve-zone-pass) beryllium, and Be-4.4 wt pct Cu and Be-5.2 wt pct Ni alloys in the temperature range of 25° to 364°C was determined using two-surface slip trace analysis, slip-step height analysis, and electron transmission microscopy. All three techniques indicated the occurrence of copious pyramidal {1 122) (1123) slip in the alloys over the entire temperature range, the amount increasing with temperature. Pyramidal slip was also indicated in the high-purity beryllium by slip trace analysis and electron transmission microscopy, but the amount was somewhat less than in the alloys. For the commercial-purity ingot crystals, only a very small number of pyramidal slip lines were observed, and these were in the immediate vicinity of the fracture surface. No pyramidal dislocations could be detected by electron transmission microscopy in this material. Dislocatransmissiontions with Burgers vectors [0001] and +(ll20) were identified by electron transmission microscopy inthe (1122) slip bands, as well as those with the j (1123) vector. This was interpreted to indicate that the edge components of the 3(1123) vector dislocations activated during c axis compression dissociate upon unloading according to the reaction i (1123) — [0001] + 3(1120) THE microstrain c axis compression of single crystals of commercial-purity ingot SR beryllium (99.6 pct), high-purity twelve-zone-pass beryllium (99.98 pct), Be-5.24 pct Ni and Be-4.37 pct Cu alloys was described in a previous paper.1 This paper covers in detail the analysis of slip traces observed on two mutually perpendicular lateral surfaces of these specimens, and a detailed description of transmission electron microscopy studies performed on foils cut from the bulk crystals after they had been deformed to fracture in the c axis compression. Observation of slip traces on single surfaces of deformed single crystals are generally insufficient to positively identify slip or twinning modes. The use of two carefully cut and oriented perpendicular surfaces can greatly aid in the positive identification and index- ing of slip traces, although even this technique may be quite inadequate if more than one type of slip system operates and if an insufficient number of traces are observed on the surfaces. The problem is greatly simplified for symmetric cases like that for c axis compression of an hep crystal such as beryllium, in which the operating slip systems are all equally inclined to the direction of the applied stress, and each slip system of a given slip mode has an equal chance of operating. For such cases, the traces of any given slip mode observed on the surfaces cut parallel to the c axis are symmetrically tilted about the c axis. It is therefore possible to quickly determine whether one or more slip modes are operating. Confirmatory evidence in support of the observations made on the external surfaces can be obtained from foils cut from the deformed crystals and examined by transmission electron microscopy. This latter technique serves to identify not only the operating slip plane but also the Burgers vector of the dislocations which participate in the slip. For this purpose, a simplified technique based upon a double tetrahedron notation is used in the present paper. The planes and directions in the hep lattice are all designated by letters rather than indices and extinction conditions are easily determined if the Burgers vector lies in the plane contributing to the diffraction. RESULTS 1) Slip Trace Analysis. The standard (0001) stereo-graphic projection of beryllium is shown in Fig. 1. The two mutually perpendicular, lateral surfaces of the compression specimen are represented by the diametrical planes AA' and BB', also referred to as surface A and surface B. For the specific case represented (a Be-5.24 pct Ni specimen deformed by c axis compression at room temperature), the A surface is tilted 5 deg to the (10i0') plane and the B surface is tilted 5 deg to the (1120) plane. Two surface trace analyses may be facilitated by examining in turn the intersection of various great circle traces of specific pyramidal planes with two surfaces and comparing the angles made with the (0001) plane with those actually observed on the two surfaces. One then identifies the slip traces by trial and error on a best-fit basis. The (1122) type planes (it was found that slip occurred on these planes) are shown plotted on the stereographic projection in Fig. 1. One obtains directly the angles between the (0001) plane and the {1122) traces by measuring the angle from the periphery to the point of intersection along the lines

Jan 1, 1969

Sulfuric Acid and Phosphate Industries at Anaconda Reduction Works. BY E L LARISON (Contribution 70-Preprint 4400 words) In 1915 a 135-ton chamber sulfuric acid plant was built at the Anaconda Reduction Works to supply acid for leaching sand tailing Sulfur dioxide was obtained by roasting copper table concentrate Soon after, the flotation process, using acid as a reagent, together with development of a leaching and electrolytic process for zinc recovery, necessitated building a second acid unit of 50 tons' per day capacity, completed in 1917 Within a few months the zinc process was so improved as to consume no added acid Mean¬while research work was in progress on production of phosphate fertilizers using sulfuric acid and Idaho phosphate rock In 1920 a pilot plant was put into operation which treated 50 tons per day of phosphate rock to produce treble superphosphate containing 45 per cent available P205 In 1922 the original plant and process for leaching of sand tailing were abandoned in favor of a combination leaching and flota¬tion process In 1923 the sand leaching plant was remodeled into a treble super¬phosphate plant designed to treat 160 tons per day of phosphate rock. In 1924 the copper flotation process was put on alkaline circuit and acid, has not since been used as a reagent Practically all the acid produced is now used for making phos¬phate products In 1931 the use of copper concentrate as a source of sulfur dioxide was abandoned and an iron concentrate, floated from the copper concentrator tail¬ing, substituted

Jan 1, 1935

By K. S. Lane

As the initial chapter in this section entitled "Role of Fluid Pressure in Rock," this attempts to fulfill an assignment for (1) summarizing the state-of-the-art and (2) illustrating engineering problems resulting from rock pore pressure. Hence, the chapter is presented in two parts: 1) Basic Concepts, and 2) Examples of Engineering Problems, including failures. Neither of these objectives has necessarily been completely met: partly from limitation of space and partly as the case histories have been restricted to the author's own field in civil engineering, and there deemphasizing those applications which will be covered in more detail by subsequent papers in this section. As with any state-of-the-art attempt, contributions of some have undoubtedly been overlooked (or omitted for brevity). Notwithstanding the foregoing limitations, it is hoped that this summary of the situation will at least: 1) Outline current concepts on effects of fluid pressure in rock-both as a pore pressure in interconnected void spaces of the intact rock and as a cleft pressure or hydraulic force acting in the joints or other discontinuities of the rock mass. 2) Indicate the often fragmentary nature of evidence presently available to support these concepts and thus encourage further research. 3) Illustrate the major significance of fluid pressure in several types of engineering problems and thus stimulate its consideration as designers encounter related situations.

Jan 1, 1970

By M. J. Salkind

In discussing the possible contribution of plastic deformation to neck growth during sintering, Wilson and Shewmon State that Lenel, Ansell, Salkind, and Early34, 35 concluded that dislocation flow is the rate-controlling process in sintering under zero external load. The experiments cited34, 35 were done with applied external loads, and the conclusions were that the same kinetics could be controlling sintering due to the force of gravity with no additional applied load. The effect of gravity on sintering behavior was previously cited.34, 31 In addition, Wilson and Shewmon argue that the conclusions of Refs. 34 and 35 are doubtful because the strain rate is proportional to stress for plastic flow. This assumption is true only for certain kinds of plastic flow such as Nabarro-Herring Microcree38 As Early, Lenel, and Ansell point out3' plastic flow, based upon dislocation movement, can exhibit strain rates which exhibit a power law relationship to the applied stress.39, 40 Wilson and Shewmon also doubt the results of Refs. 34 and 35 because the stresses in the necks are too small for x/a > 0.1. This is true, and the experiments described in Refs. 34 and 35 were done for small values of %/a. The conclusion was that plastic flow by dislocation movement was important during the early stages of sintering when x/a is small.

Jan 1, 1968

By Frederick M. Becket, Russell Franks

WideSpread experimentation has been conducted in recent years to devise a means of preventing intergranular corrosion in austenitic chromium-nickel steels of the 18 per cent chromium-8 per cent nickel type. Additions of titanium to these steels have been used commercially to overcome this dificulty, as advocated by Krupp of Germany, and the latest literature reference concerning improved details of this method is the article by Bain, Aborn and Rutherford.' Extensive work by the present authors has shown columbium also to be a particularly valuable agent for inhibiting intergranular corrosion. This paper does not deal with steels that are largely or wholly austenitic; it is limited to a discussion of the results obtained through the use of titanium and columbium in plain high-chromium steels. For many years titanium has been used commercially by some steelmakers as a cleanser, and to a lesser extent as a hardener, but then in greater percentages. The results of the work to be described disclose that titanium and columbium in proper Proportions act as effective softeners in the high-chromium steels. It is well known that chromium intensifies the hardening effect of carbon in steel; it imparts pronounced air-hardening properties within a rather wide range of chromium content; it provides increased ductility in steels of given strength and hardness; and it enhances resistance to corrosion and oxidation at elevated temperatures to a marked degree. However, in applications requiring a combination of some of these valuable properties great hardness is frequently a detriment rather than a merit. When the objective is a soft and ductile steel having optimurn corrosion resistance, the producer strives for a carbon content as low as commercially practicable. Despite these efforts some of the commercial high-chromium steels low in carbon present considerable difficulties during production, owing to their air-hardening properties, and in consequence the economic result is not favorable. A good example is found in the 5 to 6 per cent chromium steels employed chiefly in the oil-cracking industry. Sleels

Jan 1, 1934