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By S. Storchheim

MORE and more attention is being paid to the bonding of metals in their solid states. For a better understanding of this technique for joining metals and how it is affected by changes in temperature, pressure, and time at temperature and pressure, a detailed report concerning nickel to aluminum bonding has been published.' In order to broaden the knowledge accrued, some additional work concerning solid state joining of aluminum to copper and aluminum to zirconium was performed. The investigation of the Al-Cu system was considerably more extensive than the investigation of the Al-Zr system. For the A1-Cu system, not only were tensile sudies made but intermetallic penetration rate investigations also were carried out. The effect of temperature on intermetallic penetration rate for the A1-Cu system was determined at 11 tsi pressure, held 2 min. Procedure Apparatus: The hot pressing technique was the means of solid state reaction used and required the equipment depicted in Fig. 1. The following procedure was involved: The two metals to be reacted were placed in an aquadag-lubricated 18-4-1 tool steel die, 16 in. high by 1.440 in. ID, between punches of 1.366 in. diam made of the same material. A thermocouple well was located in the die body 3½ in. down from the top of the die, while another well was located centrally in the bottom punch 8½ in. from the bottom of the die. This die assembly was located in three cylindrical ceramic heating furnaces placed in tandem. Each furnace was controlled individually by a Variac power transformer. In turn, the die and furnaces were placed in a water-cooled stainless steel pot which could be evacuated. A cover, which contained a centrally located Wilson seal with an 18-4-1 1 in. diam ram running through it, was bolted on the pot. After sealing, the pot was evacuated by a roughing pump to 200 microns pressure, after which a diffusion pump was used to bring the pressure down to 5 to 15 u. At this pressure, the furnaces were turned on. As soon as they started to heat, out-gassing of the entire unit raised the pressure to 30 to 400 p. By the time the specimens were at temperature ready to be pressed, approximately 4/2 hr, the vacuum pumps had re-established the 5 to 15 u pressure. Once the desired temperature was reached, the required pressure was applied for a predetermined length of time to the 1 in. ram, through to the top punch, and to the specimen. When the time for keeping the specimen under pressure had elapsed, the pressure was released, the energizing coil current turned off, and the assembly allowed to cool. After cooling, the die was removed from the pot and the specimen was ejected. Specimen Preparation: Two different types of specimens were made for this investigation. One was for subsequent tensile testing, while the other was for determination of intermetallic alloy zone penetration into the parent metals. Tensile Bars—Commercially rolled copper pieces in. thick or zirconium sheet pieces 1/32 in. thick and 1.366 in. diam were placed between commercial 2-S aluminum rod 1 in. thick and 1.366 in. diam. This sandwich in turn was slipped into a 2-S aluminum sleeve 1.438 in. OD and 1.370 in. ID. This sleeve lined the couple up and prevented the aquadag lubricant from getting in between either the A1-Cu or Al-Zr interfaces. Immediately prior to the specimen assembly, the copper or zirconium was abraded on the flat surfaces with 320 grit silicon carbide paper, producing clean smooth surfaces. The aluminum was chemically cleaned just before assembly by: l—degrease in acetone, 2—distilled water rinse, 3—immersion for 3 min in 5 pct NaOH at 70" to 80°C, 4—distilled water rinse, 5—immersion for 2 min in 50 pct HNO3 solution at room temperature, 6—distilled water rinse, and 7—drying in a blast of gas. After the A1-Cu sandwiches were hot pressed and ejected, the specimens were machined such that the aluminum sleeve was removed, and the remaining aluminum was then threaded; the bar so produced was tested later for tensile strength. In all the instances where Al-Cu couples were tested, the specimens broke during the test at the Cu-A1 interface and never within the aluminum or copper. The ultimate tensile strength values at times showed considerable scatter for a set of given reaction conditions. Because of this, as many as three to five specimens were made for a particular set of conditions. The trend of the average tensile strengths obtained was not as conclusive as was the trend of the maximum tensile strengths, the latter values being obtained under optimum reaction conditions. Therefore, the values of ultimate tensile strength given herein are maximums.

Jan 1, 1956

By Paul M. Jr. Musgrove, Donald C. Moore

Conventional Smelting Practice Conventional copper smelting practice varies from smelter to smel¬ter, but generally consists of some or all of the following unit processes: roasting, smelting, converting, and fire refining. Roasting. Copper sulfide concentrates can be smelted directly or after an initial roasting step. Roasting is used in some smelters because roasting prior to smelting increases smelting capacity, less energy is required to melt hot roaster calcines than wet sulfide concen¬trates, roaster off gases are high in Sot concentration, 5-15% SO2, and some volatile impurities are removed from the concentrate prior to smelting. However, many smelters do not use roasters, because the problems associated with handling hot dry calcines outweigh the advantages mentioned. Concentrate roasting is performed in multiple hearth or fluid bed roasters. If the moisture is low, roasting can be performed autogenously, usually at 500-600°C. High roasting temperatures are avoided because excess oxidation of the iron compounds may lead to magnetite formation. Magnetite is detrimental to reverb operation because mag¬netite can combine with refractory minerals to form a highly viscous slag. This slag prohibits efficient matte-slag separation and leads to excessive copper losses. Also, magnetite can settle through the matte layer, deposit on the furnace bottom, and consequently reduce furnace capacity. Roasting is carried out only on sulfide concentrates prior to smelt¬ing in reverb or electric furnaces. For smelting processes, such as the flash and continuous that rely on the exothermic heat of oxidation of the sulfur minerals, roasting is not practiced. Reverberatory Smelting. The predominate copper smelting fur¬nace for the past 50 years has been the reverb. These furnaces are typically 100-120 ft long, 30-35 ft wide, and 12-15 ft high. A typical furnace layout is shown in Fig. 2. Refractory brick linings cover all internal surfaces of the furnace. Originally the flame was directed to reverberate or reflect off the furnace ceiling and melt the feed material. Current practice is to direct the flame down the furnace length to melt the concentrate. A method of charging the concentrates or calcines, generally along the side walls to minimize refractory erosion, is incorporated in the furnace design. The copper concentrates, calcines, and fluxes charged into the reverb undergo a series of complicated reactions as the temperature of the mixture increases. The reaction of the iron and copper sulfides with the oxygen in the furnace produces a molten Cu25-FeS mixture called matte. Copper smelting metallurgy is based on the fact that sulfur has a greater affinity for copper than for iron and most other common metals. Therefore, in a system containing copper, the copper will preferentially remain as a sulfide compound until all of the other metals have been oxidized. The oxidized metals combine with silica to form a silicate slag that floats on the matte and is removed from the system. Reverberatory furnace smelting chemistry can be approximated by the following chemical equations: FeS2 + O2 - FeS+ SO2 (1) The formation of FeS ensures that any copper present other than as sulfides will be reduced by the relationship: CuO2 + 2FeS + O2 - CuS + 2FeO+SO2 (2) or 2Cu +FeS - Cu2S + Fe (3) As the molten charge travels down the furnace, continued oxida¬tion of the iron minerals and sulfurization of the copper minerals occurs. When all of the copper has been converted to sulfides, the iron sulfides can then be further oxidized as: FeS + (3)2 O2 FeO + SiO2 (4) The FeO reacts with the silica added as flux in the furnace charge. A simplified equation is: FeO + SiO2 -FeO SiO2 (5) The iron silicate slag formed is skimmed from the surface at the end opposite the burners. The copper content of reverb slag is usually less than 0.6% Cu and is discarded. Matte is removed along the side wall and is taken to the converter for oxidation of the remaining sulfur and iron. The main objectives in reverberatory smelting are to produce a molten Cu2S-FeS matte containing 30-60% Cu and a throwaway slag. Production of matte permits complete conversion of all copper minerals into copper sulfides, which can migrate because of specific gravity differences, through the lighter slag layer. Also, the molten matte droplets collect the noble metals, gold and silver, as the matte settles in the furnace. The large settling area of the furnace provides enough separation time to produce a low grade slag, which can be discarded without further processing. High heat losses are associated with reverberatory smelting be¬cause of the large volume of gases sweeping through the furnace. Therefore, an outside source of heat is required to keep the smelting reaction going. Natural gas, fuel oil, or pulverized coal are used as this heat source.

Jan 1, 1985

By E. J. Roberts

The theory is developed that the tons ground through a given mesh per day in a wet ball mill is proportional to the percent plus that mesh in contact with the balls and the net power applied to the balls at this point. A grindability test is described. DURING the course of a study of the fundamentals of classification in 1937, the need for a more basic understanding of the action of a ball mill became acute. Unless one knows how classification affects grinding, one cannot hope to effectively improve on classification. The methods of evaluating grinding efficiency that depend on surface developed were studied but soon discarded for two reasons: 1. There was no apparent method which could be generally used to give a reliable figure for the actual new surface developed as a result of grinding. Subsequent papers have not changed this conclusion. 2. The practical evaluation of grinding in the main ore dressing applications was in terms of the percentage retained on a screen which passes 90 to 99 pct of the material and not in terms of surface area. The Probability Theory With the background of our experience in the field of closed-circuit grinding, together with the papers of Lennox,1 Gow,2 Gaudin,8 Fahrenwald,4 Coghill, and others, the approach of the theoretical physicist was then tried. The thought was somewhat as follows: When one grinds in a ball mill, a given expenditure of power leads either to a certain number of point to point blows per hp-hr or to a certain distance of line contact per hp-hr, depending on whether the action of the balls is considered to be cascading or rolling. It is also assumed that the balls actually come together on each blow or during the roll. Then a volume of slurry will be covered per minute which is some function of the size of the particle being considered (see fig. 1). All particles coarser than this size will be reduced through this size. This volume of slurry contains a certain weight of ore, depending on the percent solids and the density of the solids. If we fix the percent solids and the density of the solids and let w be this certain weight of ore in the volume covered, then, in mathematical terms, what we have just postulated is, w —— 8 hp (a) dt If W is the total weight of ore present in the mill, then we can write. W w/8 hp (b) W dt and if C is the cumulative percent plus the size chosen at the start of the time interval dt, w w c/dt W 8 hp x c (c) wc But wc/100 is the weight plus the size chosen which at 100 wc the close of time dt is finer than that size, and W is the decrease in the percent plus of the whole mass of ore or —dC. Then, —W dC/dt 8 hp x C. (d) In other words, the mesh tons ground through a given size per unit of time is proportional to the hp and the percent plus the mesh. A crude analogy would be to picture a 1-ft-wide steam roller going down the road at 1 ft per sec. If we place one egg on the road per square foot, one egg will be smashed per second. If we place a dozen eggs per square foot, a dozen eggs will be crushed per second. Similarly, if all the particles in w are plus the mesh, i.e., C=100, we should have a maximum rate of reduction. If only 10 pct of them are plus the mesh (C=10), we would have only one tenth the maximum rate; if only 1 pct are plus the mesh, the balls have a hard time finding anything to work on. This is where the term "probability theory" comes from. The chances of the balls crushing a particle through a given mesh depends directly on the concentration of particles coarser than this mesh in the general pulp in the mill. Giving W the units of tons and dividing equation (d) through by W, we obtain -dC hp ----- = k---— C [1] dt ton where k is a constant for any one size of particle, density of solid and moisture content of pulp. Eq 1 is the rate equation for a first order reaction and says that the rate of decrease of the percent plus a given mesh with time is directly proportional to the hp per ton applied to the body of ore and to the percent plus the mesh in the ore mass as a whole. Since it is a differential equation, it only

Jan 1, 1951

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

1.1 GENERALIZED METHODS OF SUPPRESSING THE INFLOW OF GROUND WATER DURING THE CONSTRUCTION OF SHAFTS, DRIFTS AND TUNNELS Various methods are used to prevent or minimize the inflow of ground water into underground workings during their excavation. The two most common methods include freezing the saturated rock and grouting using cement, sodium silicate, polyurethane and/or other chemicals. Each of these technologies for combatting the inflow of ground water is effective only under specific hydrogeological conditions. For example, although the freezing of saturated ground is among the more universally adopted methods, it is designed to provide only temporary protection during construction. Before the saturated rock thaws it is necessary to emplace a waterproof liner which is labor intensive, time consuming, and expensive. Consequently, freezing is used only in exceptionally complex hydrogeological conditions, namely in those cases where the water-bearing strata consist of unstable rock or the ground water has an anomalous hydrochemistry. The classical grouting of saturated rock, carried out from the surface or from the face of the underground workings for the purpose of limiting the inflow of ground water during excavation, utilizes both cement and a variety of chemical grouts. Cement grouting has been regarded as the main method of combatting the inflow of ground water in fractured rock throughout the world. In such countries as the Federal Republic of Germany (FRG), Canada, the Republic of South Africa and Great Britain, cement grouting is the main method of reducing the inflow of ground water during the excavation of mine openings. In Great Britain and the FRG, cement grouting has been used in 80% of all shaft excavations. In the Republic of South Africa cement grouting has been used in almost 100% of all the shafts that have been constructed. The installation of grout curtains into permeable water- bearing strata significantly reduces their permeability and increases the rock strength. Grouting has the greatest effect in fractured sandstones, certain well indurated shales, fractured granites, fractured quartzites, and karstic limestones or dolomites. The following principal factors must be considered when assessing the expediency of grouting rock with cement: the geometry of the network of fracture openings, the saturated hydraulic conductivity of the fractured rock, the hydraulic head acting on the water in the rock, and the chemical composition of the ground water. On the basis of the geometry of the fractures and the thickness of each hydrostratigraphic unit, the characteristics of the cement grout are selected. As a rule, the cementing of large fractured zones with high ground water velocities is carried out using inert fillers (sand, mill slag, loam, loess, crushed limestone), special types of cement, setting accelerators, and high cement concentrations in water. Calcium chloride, soda ash, sodium silicate, sodium nitrate, amino alcohols, tin bichloride, trisulphate nitrate, and lumnite are used extensively as the setting accelerators for cement grouts. In the FRG and Poland, special cement injection compounds that embody a mixture of cement and active cement metal salts, water additives and binding substances are being used for the cementing of saturated zones with rapid ground water velocities. These reagents accelerate the rate of structure-forming reactions. In spite of the wide variety of additives used for cement grouts, the effectiveness of the method in large fractures below the water table is either poor or it leads to a very large consumption of cement due to the erosion of the grout through the cracks before it hardens. At the present time, a large number of cement types and brands are being produced by various countries. These variations permit cementing to be employed in a variety of geological conditions. However, both pure cement grout and grouts with fillers constitute unstable systems with a high water loss rate. Therefore during the grouting of finely fractured rock the cement grout's premature loss of a large amount of free water causes its consistency to increase, whereupon the grout solidifies. Consequently, it inadequately penetrates into the fine fractures of the stratum. In addition the high water loss rate in the cement grout causes unreacted cement particles to remain, which greatly reduces the grout hardness and the resulting rock strength characteristics are weakened. As a result, the binding properties of the cement are utilized perhaps up to 60%. The remaining portion is left in the grout as a filler. All these conditions significantly decrease the efficiency of the isolation effort both with respect to strength and cost. In order to expand the cement grouting method, re- search is being conducted to improve the quality of cement grout, to increase its range of application with respect to permeability control, to reduce its water loss rate and to increase its capacity to withstand the erosive and corrosive effects of poor quality ground water. For example, mixtures of bentonitic clay, carboxymethyl cellulose, aerated sodium sulfide, sulphated alcohol distillery waste, nitrolignin, gip-

Jan 1, 1993

By C. L. Prokop

A laboratory investigation has been made of the effects of mud hydraulics upon the formation and erosion of mud filter cakes. The tests were conducted to simulate drilling conditions as nearly as possible. The formation of mud filter cake in a drilling well does not proceed at a uniform and unbroken rate. Instead, the rate of cake accumulation depends upon whether or not the mud is being circulated. If the mud column is quiescent, filter cake formation is a smooth function of the filtration characteristics of the system. If the mud is being circulated filter cake formation depends not only upon the filtration characteristics of the mud but also upon the erosive action of the flowing mud column Filter cakes formed during continuous mud circulation were observed to reach an equilibrium thickness after several hours' circulation. Mud circulation was maintained at a constant volumetric rate throughout each experiment. The fluid velocity at equilibrium cake thickness was dependent upon the thickness of the filter cake. Muds having exceptionally high water loss deposited thick filter cakes in spite of very high eroding velocities. The muds having good filtration characteristics deposited thin filter cakes at equilibrium circulating velocities well within tile range of those in a drilling well. It was observed that filter cakes deposited during stagnant filtration were quite difficult to erode by mud circulation. The - rate of crosion computed from the rate of filtrate accumulation after equilibrium cake thickness had been reached was in reasonable agreement with the rate of erosion obtained by direct observation. Continuous mud circulation usually caused the permeability of the filter cake to decrease with time. INTRODUCTION Many of the difficulties encountered during tile drilling of a well have been attributed to the loss of water from the mud and the attendant deposition of solids upon the walls of the hole. Past experience has shown that a reduction of the filtration rate of the drilling fluid eliminates or greatly reduces these difficulties. Definite filtration requirements, however, are hard to establish for a given set of conditions. This is due. in part, to the fact that the usual filtration test performed upon mud doe? not simulate well conditions as closely as desirable. The filtration characteristics of a mud are customarily determined by means of the standard low-pressure API wall-building tester.' In this instrument a filter cake is deposited upon a horizontal bed under a pressure differential of 100 psi. The rnud is quiescent during the filtration period. In actual practice. mud filtration occurs within a well under quite different conditions. One of the major differences is that mud flows upward across the filter bed as the filter cake forms. This undoubtedly produces a change in the filter cake which cannot be reflected in the results of the API test. The laboratory work described in this paper had as its primary objective a better understanding of the influence of mud circulation upon the thickness and ,characteristics of the filter cakes deposited under conditions similar to those existing in a drilling well. ANALYSIS OF PROBLEM Once a permeable formation is penetrated by the bit, filtrate from the mud flows into the formation. 'he mud solids plaster against the walls of the hole, forming a filter cake. If the mud column is stagnant, that is, if it is not being circulated. the filter cake will increase in thickness until the hole is filled. Prior to the time that the hole is filled, the thickness of filter cake existing at any given time will be a function of the filtration characteristics of the mud, the temperature, and the pressure differential. The effects of these variables have been investigated in the past for both flat bed filtration2'3 and for radial filtration.' When the mud is circulated in a hole in which a filter cake i. being deposited. some of the solids that would ordinarily deposit in the filter cake will be carried away by the eroding action of the mud. This will limit. filter cake thickness. Some work has been done to determine the effect of flow upon the filtration rate in a circulating mud system' but little work has been done upon the factors which determine the filter cake thickness existing in a circulating system. On first sight it would appear that the major factors controlling filter cake formation in a circulating system should be: 1. The rate of deposition of solids from the mud. 2. The erosive force that the flowing mud exerts upon the filter cake. 'A. The erodabilitv of the filter cake. 4. Any change in filter cake characteristics attributable to the scouring action of the mud. The rate at which solids are deposited from the mud will be controlled to a large degree by the filtration characteristics of the mud, the pressure differential. the temperature under

Jan 1, 1952

By Paul A. Taylor

MUCH information concerning pegmatites which was thought to be true a few years ago has been proved false, and what is now actually known about some pegmatites is not true of many others. The erratic and seemingly unpredictable structure and variable composition of this class of mineral deposits has been widely emphasized. Even parts of the same pegmatite body exhibit marked differences in texture, mineral composition, width, and attitude. Constructive geological thinking in respect to pegmatites now aims to establish general laws rather than to stress the confusing diversity of features having no special economic significance. Substantial progress has been made in classifying different types of these deposits according to general features, internal structure, mineralogy, and origin. In some cases it has even been possible to block out tonnage reserves in advance of mining. It is still easy, however, to make highly erroneous predictions after a preliminary examination of a pegmatite prospect. Pegmatites are important to the economic well being of the country and to its military security. They furnish virtually all the feldspar, strategic mica, beryl, columbium, tantalum, and caesium utilized in the United States, as well as sundry other minerals and significant amounts of lithium and rare earth minerals and gems. With the exception of vermiculite, occasional ilmenite-rutile, and perhaps soda-lime feldspar and garnet deposits, basic pegmatites are of little economic importance. Consequently in this paper, as in common parlance, the term pegmatite generally relates to coarsegrained acidic rocks or what is aptly called giant granite. Available data indicating the size and importance of the production and trade in specified pegmatite minerals are summarized in Table I. Geological Features Much of the latest thinking on the economic geology of pegmatites is now available in a 115-page monograph' by a group of experts who participated with geologists of the Federal Geological Survey in the widespread wartime investigations. Doubtless the most significant feature of the monograph is indicated by the title, The Internal Structure of Pegmatites, but it also contains a vast amount of other new information and includes the assimilated concepts of many earlier writers, whose works are given in a comprehensive list of references. The shape, size, attitude, and continuity of many pegmatite bodies is controlled by the structure of the older rocks in which they occur. If the older rocks are easily penetrated, e.g., biotite schist, most of the pegmatites in a given district will be found outside the parent granite mass as exterior pegmatites. Marginal pegmatites are more prevalent if the older rocks are massive, unsheared, and sparingly jointed. Networks of pegmatites are abundant in highly-jointed rocks. In strongly foliated schists the bodies are usually lenticular, whereas in highly-folded areas they assume tear drop, pipe or pod-like, bow-shaped, or sinuous forms. Jahns2 recognizes five major shape classes: l—dikes, sills, pipes, and elongate pods; 2— dikes, sills, pipes, and pods with bends, protuberances, or other irregularities; 3-—trough-or scoop-shaped bodies with or without complicating branches; 4—bodies with the form of an inverted trough or scoop; and 5—other bodies, including combinations of the above and miscellaneous shapes. Many pegmatite deposits are scarcely big enough to be recognizable as such. Most of them, in fact, are small tabular deposits less than 4 in. wide and usually without economic concentrations of minerals. On the other hand, some pegmatites are more than a mile long and over 500 ft wide. The ratios of length to breadth range from 1 : 1 to 1 : 100. Although the vertical dimension bears no invariable relationship to strike length, tabular deposits or large lenses are often symmetrical enough to show nearly as much continuity down dip as in their horizontal extension, and some pipes or pods are amazingly persistent in the vertical plane. Small pegmatites often string along a fairly definite trend line; in a given district major bodies may lie roughly parallel, and where only a few of them do not, the erratically disposed bodies generally differ in composition from those conforming to the regular pattern. This does not apply, however, in all districts. Characteristically, pegmatite veins pinch and swell or split into branches. When they pinch out entirely it is often possible to find a new body by prospecting the extension of the strike or dip, but the chances of finding a hidden deposit are ordinarily too uncertain to justify much subsurface prospecting. Diamond drilling may yield valuable information as to the continuity of known deposits whose upper portions are well-exposed. Some deposits, in fact, can be proved up for hundreds of feet by surface trenching and then intersected by drill holes at various depths like any other vein-like deposit. Others twist and branch, apparently defying all efforts to explore them short of actual mining.

Jan 1, 1954

By Samuel H. Dolbear

THE purpose of a technical report is to record facts, usually collected by investigation, and to interpret these facts in understandable language. The audience may range from a small shareholder without technical knowledge, to a highly trained engineer or geologist. If the client is a mining company with technically trained executives, the report writer's problem is relatively simple. The writer will then be appraised not only for his conclusions, but for clarity of language and organization of the report. If the writing is bad, the construction careless, and there is a failure to clearly convey to the reader the facts and the author's conclusions, the report is a failure and the writer may have damaged his professional character. Spoken words die quickly, written words may constitute a permanent record and if they are badly composed they may rise up to damn the author. "A good measure of an author's understanding of his subject is his ability to express it clearly in plain words." Good English Teachers in the fields of mineral technology have frequently complained that even in post-graduate groups there is an appalling indifference to their appeals for good English. Some have even noted a student's belief that the use of refined English is effeminate. If those with such immature beliefs could measure the pay-check damage arising from the use of "sloppy" language, they would realize that precision and refinement in English may be quite as important as technical accuracy. When the reader audience is without technical knowledge simplicity in treatment becomes especially important. If one is engaged in consulting work, in government service, or in any field where reports have public distribution, the language employed should be technically adequate but simple enough to be understood by non-technical readers. For example, one may use the term "visual" in place of "megascopic." Technical language can be so obscure that it cannot be understood even by highly-trained students. In the March-April (Vol. 47-No. 2) issue of Economic Geology, Nicholas Vanserg ridicules these extremes and quotes various paragraphs from published material, such for example: "However, lattice orientation unaccompanied by cognate dimensional orientation can never be attributable to growth from an isotropic blastetrix." "The temperature declines because of cessation of the exothermic chemical and mechanical equilibriopetal processes." These he calls "good geologese" and they are calculated not only to baffle the reader but to impress upon him the erudite character of the author. Revision In some cases, difficulty arises from the fact that the writer is too close to the subject and unconsciously assumes that his reader is equally familiar with the background of the report. It is difficult for the writer to regard his work objectively and to determine to what extent it is likely to be understood. Every important manuscript will gain in clarity if the author will have it reviewed by an informed reader. But the writer must not be oversensitive to criticism and should not treat his composition as perfect and beyond the possibility of improvement. The first draft of a report always requires revision, regardless of the care used or the ability of the writer. Three or four revisions are not uncommon. The first draft usually requires expansion in places, the deletion of non-essential material, and language changes to promote clarity of expression. This should be done by the author after a lapse of time, even if only overnight, in which his mind has been occupied on some other subject. Possible improvements are always more visible. The manuscript should be passed on to another reader for further suggestions. Organization of Material The engineer should study available reports and library references before going into the field. If the previous reports have been responsibly done and can be accepted as correct, then much field time can be saved. It is, of course, customary to make some on-the-ground checks to confirm earlier reports, particularly those relating to ore reserve which may have undergone changes. Report writing requires time and expense, but nevertheless, the basic reasons for conclusions should be presented even in the case of a worthless property, for it may prevent a duplication of the work. If the mine examined is obviously of no further interest, no useful purpose can be, served by preparing a report in detail. In one case an engineer travelled all the way to South America only to find that the mine had been grossly misrepresented and was valueless. His cabled report "Nonsense" is a case of over-simplification, but it served his company's purpose. The first step in report organization should be the selection of subjects. This should be done at the mine, and the data collected for each subject should be reviewed in considerable detail on the ground. Otherwise one may find that he has failed to collect some essential details not readily obtained after he has returned to headquarters. If the property to be described is undeveloped, then many of the subject titles are automatically eliminated. Usually no useful purpose is served by an attempt to calculate the cost of production under such circumstances, although the cost of exploration

Jan 1, 1952

By E. Douglas Sethness

The uranium industry is booming. In Texas alone, there are about 22 different companies with active exploration programs. Twelve solution mines have been permitted; three surface mines have been authorized; and two mills are currently in operation. However, the industry also has a problem, and that is the disposal of radioactive wastes. Over the past several years, stories concerning nuclear wastes have appeared frequently in the news. One of the most frequently cited cases occurred in Grand Junction, Colorado. In 1966, after ten years of investigations, the U. S. Public Health Service (PHS) discovered that tailings from a uranium mill were being used as fill material and aggregate for local construction purposes. It was estimated that between 150,000 and 200,000 tons of material had been removed and used under streets, driveways, swimming pools, and sewer lines. In addition, tailings had been used under concrete slabs and around foundations of occupiable structures. Further studies prompted the Surgeon General to warn that the risk of leukemia and lung cancer could be doubled at the measured radiation levels. More recently, the L. B. Foster Company discovered that its building site in Washington, West Virginia, was radioactive. While digging a foundation, the ground erupted and a ball of fire 30 feet high shot out. Evidently, the dirt was laced with radioactive thorium and zirconium, a potentially explosive mixture contained in a Nigerian sand which had been used by the previous site owners in the manufacture of nuclear fuel rods. Just this month we have read about legal suits to stop exploration for a nuclear waste disposal site in Randall County, Texas. The U. S. Department of Energy is trying to locate a deep underground nuclear waste depository for final burial of over 76 million gallons of high-level wastes. The problem is acute, the wastes are accumulating at a rate of about 300,000 gallons per year. Nor do these numbers include the spent fuel elements from nuclear power plants that are in temporary storage facilities. Fortunately, public awareness of these and other related issues is high. Unfortunately, the differences in the waste products from the nuclear fuel cycle are not always apparent to the general public. There are two distinct types of radioactive wastes: "high-level", which consist of spent fuel or wastes from the reprocessing of spent fuel; and "low-level", which, in general, are by-product wastes. There are numerous non-technical definitions that can be applied to help the layman differentiate between high-level and low-level wastes. For this latter purpose, it is best to think of them in terms of what we can see and feel. In general, high-level wastes are physically hot and can cause acute radiation sickness in a short period of time. Low-level wastes are not hot, but may cause chronic health effects after long exposure. The wastes which we are concerned with in the uranium mining and milling industry are low-level wastes. As recently as ten years ago, there were very few controls or regulations governing tailings disposal methods. At the same time, mine reclamation was not enforced through either state or Federal laws and the long-term viability of abandoned tailings ponds was not assured. The regulatory climate has changed significantly in the last decade, however. The low-level radioactive wastes generated by uranium mining and milling are generally contained in a tailings pond. Approximately 85-97% of the total radioactivity contained in uranium ore is present in the mill waste that goes to such tailings ponds. The isotope Radium-226 is probably the most potentially harmful radioactive parameter in the ponds. Radium emits gamma radiation and is also an alpha particle emitter. Because gamma radiation is very penetrating, it presents a potential health problem when a source is located external to the body. Gamma radiation will go through the body, causing damage to each cell encountered on the way. Although alpha particles have very little penetration capability, they can cause extensive cell damage. For this reason, alpha particles are a problem after inhalation or ingestion. Radium creates a health hazard by both of these mechanisms. Radium decays to radon gas which can be inhaled and serve as an alpha particle emitter. Additionally, radium is very soluble and readily enters the natural hydrologic cycle if allowed to leach from a tailings pond. With a half-life of 1620 years, radium has plenty of time to be taken into the food chain and end up in our bodies, emitting alpha particles. Because the potential health problems are better understood today than ten years ago, and because the Nuclear Regulatory Commission (NRC) has developed increasingly stringent government regulations, the uranium mining industry applies a high level of technology to the disposal of nuclear wastes. In most cases, low-level radioactive wastes are disposed of at or near the site where they are produced. There are six commercial burial grounds for low-level wastes, but it would not be economical to ship all mine or milling wastes to these sites. The on-site disposal methods most often used are ponding

Jan 1, 1979

By G. M. Schwartz, D. M. Davidson

THE Duluth gabbro outcrops containing sulphides of copper, nickel, and iron are located on both sides of State Highway No. 1 an airline distance of 8.5 miles southeast of Ely in northeastern Minnesota. The region of known sulphide occurrences includes parts of sections 5, T. 61 N., R. 11 W., and parts of sections 25, 26, 32, 33, and 34, T. 62 N., R. 11 W. These sections, given in Fig. 1, are all in Lake County, Minnesota. Part of the area, which lies entirely within the Superior National Forest, is shown on the topographic map of the Ely quadrangle. The original discovery was made in 1948 when a small pit was opened in weathered gabbro rubble for use on a forest access road. A shear zone had caused unusual decomposition in this glaciated area, and the resulting copper stain was noted by Fred S. Childers, Sr., an Ely prospector, who began searching the outcrops along the base of the intrusive. He was joined in further exploration by Roger V. Whiteside of Duluth. In the summer of 1951 a small diamond drill was moved into the area and a hole 188 ft deep was drilled, passing through 11 ft of glacial drift into sulphide-bearing gabbro. This paper is a preliminary report on the geology of the newly discovered ore. The Duluth gabbro is one of the largest known basic intrusives and may be defined as a lopolith.1 It extends northeastward from the city of Duluth as a great crescent-shaped mass that intersects the shore of Lake Superior again near Hovland, 130 miles to the northeast, see Fig. 2. The distance around the outside of the crescent is nearly 170 miles. The form of the intrusive is simple at Duluth where it ends abruptly north of the St. Louis River; at the east end, however, the gabbro splits into two elongated, sill-like masses separated mainly by lava flows and characterized by minor irregularities. The outcrop reaches a maximum width in the central part where it is about 30 miles across, and a maximum thickness of about 50,000 ft. It may be significant that the sulphides occur at the base of the thickest part. The lopolith has segregated into rock types ranging from peridotite to granite. The most abundant types are olivine gabbro, gabbro, troctolite, anorthosite, and granite. Of lesser importance quantitatively are peridotite, norite, pyroxenite, magnetite gabbro, and titaniferous magnetite. Grout estimates that two-thirds of the gabbro at Duluth is olivine gabbro. Variations in the percentages of plagioclase, augite, olivine, and magnetite-ilmenite constitute the only essential differences found among the basic rock types. The predominant mineral is plagioclase, mainly labradorite. Plagioclase and olivine seem to have crystallized early, and the olivine rich rocks, usually troctolite, are found in the lower part. Segregations of titaniferous magnetite are abundant near the base of the gabbro along the eastern part and also occur far above the base. These have recently been described in detail by Grout' Near the top, segregation has produced a gradation to granite, or "red rock," as it is known locally. This consists of quartz, red feldspar, and hornblende. The red rock forms a. zone with a maximum width of nearly 5 miles but is quantitatively unimportant from Duluth northward for 35 miles. In Cook county, where the gabbro splits, each of the two sill-like masses has a red rock top somewhat thicker in proportion to the gabbro below than in the main central mass. The intrusive ranges from coarse to medium in grain size and from granitoid to diabasic in texture. Throughout much of the Duluth gabbro in Minnesota banding and foliation are well developed, as Grout has emphasized! The bands are mainly a result of variation in the percentage of minerals, as in troctolite with alternating bands high in olivine and in plagioclase. A few bands may consist largely of one mineral, as is true of some segregations of magnetite. Many of the banded rocks show a clearly developed parallelism of platy plagioclase crystals, and both banding and foliation are believed to conform to the floor of the lopolith. Throughout its extent in Minnesota the Duluth gabbro dips east and south toward Lake Superior. It is generally believed to extend beneath Lake Superior and is found as a smaller mass exposed along the north side of the Gogebic district in Wisconsin and Michigan. The dip at and near the base ranges along most of its length from 20 to 40°, but at places the internal banding dips even more steeply. The dip of the upper part is much less, and if it is assumed that the flows along the north shore of Lake Superior are a dependable indication, it does not exceed 15º. The formations shown in Table I which are intruded by the gabbro range from Keewatin to Middle Keweenawan in age. They present a significant picture. At the top, the gabbro and its accompanying

Jan 1, 1952

By L. D. Hall, D. R. Mash

The paper presents the results of experiments on the anelastic. behavior of gold, as manifested by grain boundary relaxation. Two grain boundary internal friction peaks are found for 99.9998 pct Au. It is found that the peaks are associated with primary and secondary recrystallization. However, the existence of two discrete peaks cannot be explained on the basis of grain size and shape alone. It is suggested that grain boundary stability, as determined by orientation, plays a role in the observed effects. EVIDENCE for the viscous behavior of grain boundaries in metals has been presented in recent years by several investigators, based upon studies of various anelastic effects, especially internal friction. KG1 has contributed greatly to this field, having put forward a coherent body of evidence for stress relaxation by the viscous intercrystalline flow mechanism. In this connection, he has made extensive use of pure aluminum (99.991 pct) as the test material, although he has also studied other metals and alloys, including pure iron (Puron).² Rotherham, Smith, and Greenough³ have studied the internal friction of pure tin, interpreting their results in a manner similar to that of KG. In view of the importance of such studies in shedding light upon the fundamental structure and behavior of the grain boundaries in pure metals, it appears that the use of a very pure test material which is inert to its environment should provide useful information on anelastic properties and the source of such behavior in pure metals. The present work was carried out on spectrograph-ically pure, 99.9998 pct Au, free of all impurities except for a trace of silver, estimated to be present to the extent of about 0.0002 pct. The term "pure gold" will hereafter refer to this very pure material. Gold of commercial purity, 99.98 pct, was also studied to observe the effects of small amounts of impurities. A pure gold "single crystal" specimen was also tested for comparison. The variation of the internal friction and rigidity modulus as a function of temperature was determined by means of a torsion pendulum apparatus employing extremely low stress amplitudes and a frequency of vibration of the order of 1 cycle per sec. A 12 in. length of 0.031 in. (20 gage) gold wire formed the suspension element. The apparatus was similar to that described by Ke.l The test procedure and the basic requirements to be met for obtaining useful experimental data by this method have been given elsewhere.1,2 It should be made clear that in all of the experiments to be described, the internal friction and rigidity were independent of the amplitude of torsional vibration. The semilog plot of amplitude of vibration vs ordinal number of vibration was a straight line. This was carefully verified for each internal friction measurement. The linear variation shows that the internal friction was independent of stress; i.e., that the specimens were not being cold-worked during testing. The reproducibility of the internal friction curves, which were obtained by cyclic heating and cooling, indicates that the gold was unaffected by its environment during the tests. The measure of internal friction adopted in the present study is the conventional "logarithmic decrement," defined as follows: log. dec. = l/n In A0/An [I] where n is the number of cycles or vibrations; A,, the initial amplitude of vibration; and An, the amplitude after the nth cycle. When the logarithmic decrement is small, the shear modulus, G, of the wire is proportional to the square of the frequency of vibration provided the length and radius of the wire are kept constant. A plot of frequency squared vs temperature gives the following ratio:' This expresses the fraction of the stress which has not been relaxed at a given temperature. Gr and Gv are the relaxed and unrelaxed moduli, respectively. The frequency of vibration in the polycrys-talline specimen is fp, and the frequency of vibration of a single crystal is f8. This latter quantity is obtained simply by extrapolating the linear, low temperature portion of the curve of frequency squared vs temperature for the polycrystalline specimens. The theory of viscous grain boundary stress relaxation as demonstrated by the anelastic behavior of metals has been discussed in detail by Zener4 and need not be reproduced here. Experimental Results Initial measurements of the internal friction of pure gold were carried out on specimens which had been drawn with no intermediate annealing, resulting in a material which had undergone approximately 99 pct reduction of area in final processing. Annealing was then carried out at successively higher temperatures starting at 400°F for 1 hr and proceeding in this manner to as high as 1600°F in 100°F intervals. After each annealing treatment an internal friction and rigidity vs temperature curve was obtained over the range from room temperature to the particular annealing temperature. The resulting internal friction curves did not exhibit well defined maxima (peaks), but rather several fairly flat

Jan 1, 1954

By S. L. Bessinger

Introduction In the design and specification of a ventilation system for an underground mine, it is necessary to make reasonably accurate estimates of the pressure losses in the various airways of the mine. These estimates can be made with little difficulty for open airways with simple geometric cross-sections, such as those cut by continuous miners or tunnel-boring machines. The situation is much different on a longwall face, where the airway's complex geometric cross section and the presence in the airway of obstructing equipment having a variety of shapes make it difficult, if not impossible, to estimate pressure loss using traditional methods of calculation. Head losses in mine entries are calculated using Atkinson's Equation. [22H= KPLQ (English) H= KP 3Q (SO (1) 5.2AA] where H = pressure loss, in. of H2O (Pa); K = friction factor, lbf•min2/ft4 (kg/m3); P = perimeter, ft (m); L = airway length, ft (m); Q = airflow quantity, ft3/min (m3/sec); and A = flow cross-sectional area, ft2 (m2) In this equation, the friction factor, K, is an empirical constant that describes the aerodynamic roughness of the airway. Typically, the K-factor for a given airway is determined by measuring the factors H, P, L, Q and A in Equation (1) and calculating K. Tables of friction factors calculated in this way are found in textbooks and handbooks that deal with mine ventilation analysis. Unfortunately, very few K-factors have been measured on longwall faces, and the accuracies of those that have been measured are entirely site specific, because of the wide variety of equipment found on longwalls. The development of a technique for prediction without mine-site measurements of the friction factor for any longwall face, regardless of its configuration, will thus be very useful in the design of ventilation systems for mines in which longwall mining is practiced. Calculation of pressure losses using Atkinson's Equation (1) and empirically determined K-factors provides accurate and useful approximations in cases where the airways have relatively simple cross sections. However, a careful analysis using the principles of fluid mechanics shows that such calculations are based on two assumptions that are not strictly correct when there are obstructions in the airway. The first assumption is that the air velocity distribution in the cross section, particularly around the perimeter, is uniform. This assumption results from the fact that the tabulated K-factor values found in the literature are based on field measurements with uniform conditions. Such uniformity does not exist in longwall airflows. The second common assumption is that the K-factor, and corresponding head loss, is independent of the Reynolds Number (NR) for a given flow. In fact, this assumption is not strictly correct, and is particularly erroneous in the case of irregular protuberances into the airflow, such as those found on a longwall face. The errors arising from the assumptions may be avoided by using K-factors calculated by a newly devised method, described below, which takes into account the fundamental principles of aerodynamic drag analysis. This new technique has two advantages: first, it is flexible enough to model any longwall, regardless of equipment configuration; second, it employs terminology and equations familiar to those who perform mine ventilation analysis, using K-factors, for which ventilation engineers have an intuitive understanding, rather than drag coefficients. To provide guidance for development of a longwall drag model, data were taken on two modern longwalls operating in substantially different conditions. Pressure measurements at Mine B were made with 200-foot (61-m) sections of 1/8-in. (3-mm) diameter plastic tubing, attached to a Dwyer Magnehelic gauge. Pressure drops were measured in 200-foot increments down the face, and summed to give the drop for the entire face length. This technique was found to produce small, repeated errors because of the number of segments required to span the longwall. At Mine A this problem was avoided by using a single, continuous, plastic tube for the entire face length. The psychrometric properties of the air were measured for both Mines A and B. A calibrated, standard vane-anemometer was used to measure the airflow on both faces. Finally, numerous dimensions were measured on both faces, and face profile drawings were obtained to allow detailed evaluation of the face equipment geometry. From this information, accurate evaluations of the average wetted perimeter and average area of the longwall face airways were made. Since the airflow is not confined to inside the powered supports at all points along the face, a quadratically weighted average of the airflows measured at various stations along the face was calculated: [n2Qavg =Qi Ii / It(2)i=1] where [Q, avg = average airflow for analytical purposes, ft3/min (m3/sec); Q= airflow at station i, ft3/min (m3/sec); 1= length of segment represented by Q, ft (m); h= length of longwall face, ft (m): and n= number of quantity measurement stations.] The quadratic weighting scheme was chosen because the

Jan 1, 1992

By O. N. Carlson, S. A. Bradford

Discontinuous yielding in tensile tests was observed in V-O alloys in the temperature ranges of 150° to 175°C and also 350° to 400°C. The magnitude and intensity of the serrations were found to vary considerably with oxygen content. Maxima were observed in tensile and yield strengths and in the strain-hardening coefficient at the higher temperature only. The strain rate sensitivity was observed to be negative between 150° and 400°C. THIS investigation was undertaken to study the effect of oxygen on the tensile properties of iodide vanadium in the temperature range of 25o to 450°C. Brown1 observed an increase in strength between room temperature and 400°C in vanadium metal, and found that oxygen and nitrogen had a rather pronounced effect on the strength and ductility. A maximum in the tensile strength was observed by Rostoker et al.2 near 300oC and by Pugh3 around 450°C for calcium-reduced vanadium. Pugh also found a maximum in the yield strength and in the strain-hardening exponent, and minima in the elongation and strain rate sensitivity at the same temperature. Eustice and Carlson4 reported the appearance of serrations in the stress-strain curves between 140° and 180°C in iodide vanadium containing 600 ppm O. These anomalies in the mechanical properties indicate that strain aging occurs in vanadium, but the impurity or impurities responsible for the above-mentioned effects have not been identified. The phenomenon of strain aging is usually characterized by the return of the yield point after interruption of a strength test. In the temperature range where strain aging occurs, the yield and tensile strengths attain maximum values, elongation and strain rate sensitivity exhibit minima, and discontinuous yielding is generally observed in the stress-strain curve. Cottrell5, 6 has postulated that strain aging is due to the migration of solute atoms to dislocation sites to produce locking after the dislocations have broken free from their impurity atmospheres during the initial yielding. At the strain-aging temperature the process is a dynamic one in which the solute impurity atoms diffuse to the vicinity of the moving disloca- tion producing "locking" which gives rise to maxima in the tensile strength and serrations in the elongation curves. Cottrel17 has noted that discontinuous yielding in iron occurs when the diffusion coefficient of nitrogen, D, and the strain rate, i, are related by D = 10-9 €. EXPERTMENTAL PROCEDURE The vanadium metal employed in this study was prepared by the iodide refining process as described by Carlson and owen.8 A representative analysis of the vanadium used in this investigation was: 150 ppm O, <5 ppm N, <1 ppm H, 150 ppm C, 150 ppm Fe, 70 ppm Cr, <50 ppm Si, 30ppm Cu, 20 ppm Ni, <20 ppm Ca, <20 ppm Mg and <20 ppm Ti. Alloys containing from 200 to 1800 ppm O, all of which lie in the solid solution range of the V-O system, were prepared by arc melting vanadium together with portions of a high-oxygen master alloy. The master alloy was prepared by tamping pure V2O5 into holes drilled in a vanadium ingot and arc melting this five or six times in an inert gas atmosphere, inverting the button between each melting step. The oxygen content of the master alloy was then determined by vacuum fusion analysis. Vanadium containing less than 150 ppm O was prepared in the following manner. A bar of iodide vanadium was deoxidized by sealing it in a tantalum crucible with a few grams of high-purity calcium. This was held at 1100°C for 4 days to allow time for the oxygen to diffuse to the surface and to react with the calcium vapors. The calcium oxide product was later dissolved from the surface of the bar with dilute acetic acid. In this way vanadium containing from 20 to 50 ppm O was prepared. Sample Preparation. The are-melted ingots were cold swaged into 3/16-in. diam rods and these were machined into cylindrical tensile specimens with a reduced section of 1.00-in. length and 0.120-in. diam. The test specimens were annealed for 4 hr at 900°C in a dynamic vacuum of mm of Hg to remove hydrogen from the metal. This recrystal-lization treatment produced a uniformly fine-grained structure with a mean grain size of approximately 0.06-mm diam. The oxygen contents reported in this paper were determined by a vacuum fusion analysis of the tensile specimens after testing. Analyses for other interstitial or metallic impurities showed no significant changes from that of the original material. Tension Tests. Tension tests were performed on a screw-driven tensile machine at a constant cross-head speed of 0.01 in. per min. Tests at elevated temperatures were carried out by heating the

Jan 1, 1962

By O. Mellgren, A. M. Gaudin, P. L. De Bruyn

The adsorption density of ethyl xanthate on pyrite was determined as a function of xanthate concentration. Surface preparation of the mineral appears to have asafunctionsome effect on the subsequent adsorption process, A monolayer of xanthate on the surface is exceeded only in presence of oxygen. The effect of OH- , HS- (and x and CN- S=)and on the amount of xanthate adsorbed was investigated. Competition between OH- and X- (xanthate) ions for specific adsorption sites is indicated over a wide pH range. IN the flotation of sulfide ores, xanthates are most commonly used to prepare the surface of the mineral to be floated so that attachment to air takes place. The quantity of agent required to make the mineral hydrophobic is usually very small, of the order of 0.1 to 0.25 lb per ton of mineral. Details of the mechanism of pyrite collection are for the most part unsettled. Adsorption of collector has long been believed to involve an ion exchange mechanism as demonstrated for galena&apos; and for chalcocite.2 In the work on chal-cocite it was also demonstrated that a film of xanthate radicals unleachable in solvents that dissolve alkali xanthates, copper xanthate, or dixanthogen was formed at the surface of the mineral. The unleachable product increased with increasing addition of xanthate up to a maximum corresponding to an oriented monolayer of xanthate radicals. Pyrite is extremely floatable with xanthate if its surface is fresh.9 ut the floatability decreases rapidly as oxide coatings increase in abundance. Pyrite shows zero contact angle when in contact with ethyl xanthate solution at pH higher than about 10.5;4 at neutrality, a contact angle of 60" is obtained at a reagent concentration of 25 mg per liter. Alkali sulfides and cyanides are pyrite depressants. In this study of pyrite collection the writers have sought to relate measured xanthate adsorption to the method used in preparing pyrite, to the presence or absence of oxygen, to concentration of hydroxyl, hydrosulfide, sulfide, and cyanide ions. The principal experimental tool has been radioanalysis," " using xanthatcx marked with sulfur 35. Experimental Materials Pyrite: Unlike most sulfides, pyrite is a poly-sulfide. The structure given by Bragg7 resembles that of sodium chloride, the iron atoms corresponding to the position of sodium and pairs of sulfur atoms corresponding to the position of chlorine. The edge of the unit cell in pyrite is 5.40 A and in halite 5.63 A. The S-S distance in pyrite is 2.10 A; the Fe-S distance, 3.50 A: and the Fe-Fe distance, 3.82 A. Natural pyrite from Park City, Utah, was used in this investigation. Pyrite 1 was obtained by hand picking pure crystals. Pyrite 2 and Pyrite 3 were obtained from finer textured crystalline material containing inclusions of silicates. The same cleaning technique was utilized for the preparation of Pyrite 2 and Pyrite 3, whereas a different cleaning technique was used for Pyrite 1. Pyrite 1 was prepared as follows: The crystals were ground in a porcelain ball mill and the 200/400 mesh fraction was separated by wet screening with distilled water, followed by washing for 1 hr with deoxygenated distilled water acidified with sulfuric acid to pH 1.5. The acid was removed by rinsing with deoxygenated distilled water on a filter until a pH of 6.0 was reached in the effluent. This filtration was carried out under nitrogen. The sample was then dried in a desiccator under nitrogen. The period of time for which this pyrite sample was in contact with water containing oxygen was about 4 hr. The specific surface as determined by the BET gas adsorption method was 582 cm2 per g. Final material assayed 53.12 pct sulfur and 46.5 pct iron (theoretical, for FeS,: S, 53.45 pct; Fe, 46.55 pct). After crushing, Pyrite 2 and Pyrite 3 were washed with 1 M HCl. rinsed, and fed to a laboratory shakinq table to remove the small amount of silicates. The concentrate obtained was ground in a laboratory steel ball mill. The 200/400 mesh fraction was separated by classification in a Richards hindered settling tube. This fraction was then given a final wash with 0.1 M HCl and deoxygenated water was filtered through the sample. The final effluent showed a conductivity equivalent to that of a solution having a salt concentration of 0.3 ppm. Aqueous hydrogen sulfide solution was then added to the sampln (about 100 ml saturated H,S solution to about 1000 g pyrite under a few hundred milliliters of water) which was stored wet under nitrogen. The sample stored in this manner showed no indication of formation of iron oxides, whereas iron oxides appeared

Jan 1, 1957

By F. W. Glaser

Metal-carbon-boron powder mixtures were hot pressed and the resulting specimens were studied by X-ray diffraction. It was found that regardless of the starting combination of the metal, carbon, or boron powders, a metal boride phase was always the major component in these samples. In the absence of carbon the boride phase formed on hot pressing depended only on the amount of boron present. Two new phases of the system Ti-B were found. They are Ti2B and Ti2B5. The existence of a controversial face-centered cubic phase of formula TiB was confirmed. Electrical resistivities were measured for various boride phases. It was found that the diborides are generally better conductors than the monoborides of the same metal. THE carbides and borides of the transition elements have very high melting points, in the range 2500° to 4000°C, and are therefore of interest as high temperature materials. The literature on the stability or chemical reactivity of these carbides and borides is very scarce. Various investigators&apos;-" have demonstrated a relative instability of certain carbide phases in the presence of boron or boron-containing substances. In a recent publication, Glaserl demonstrated the stability of zirconium-boride (ZrB,) in the presence of carbon at temperatures in excess of approximately 2900°C, while during a preliminary investigation of boride phases, Steinitz&apos; concluded that the diborides are stable in the presence of carbon while the monoborides of the fourth and fifth group are not, forming diborides plus carbides instead. Nelson, Willmore, and Womeldorph" have elaborated on the reaction B,C + 2TiC = 2TiB, + 3C, which was known to occur because of a relative instability of B,C and the great tendency towards TiB, formation at relatively low temperatures (approximately 1200°C). A similar study, involving as starting materials TiO, and B,C and resulting in TiB,, was recently described by Honak4, who observed the beginning of an exothermic reaction of a Ti0,-B,C powder mixture, which, when preheated in a hydrogen atmosphere to approximately 950°C, was carried to about 1600 °C by the heat of reaction. To shed more light on reactions of this type (Metal-C-B), the final product apparently always resulting in a boride phase at the expense of a carbide phase," a systematic investigation was started * Boride phases of various metals, as reported to date, are listed in Table I. and the following is an account of some of the results that were obtained. Materials, Preparation of Samples, Testing Methods The raw materials employed for this work consisted of various carbide, boride, and metal powders. as well as of boron and graphite powders. In cases where commercial grades of carbides were considered unsuitable because of low purity or excessive amounts of graphitic carbon, such carbide powders were prepared by this laboratory. The procedure for the preparation of carbide powders (zirconium carbide, titanium carbide, tantalum carbide, and niobium carbide) consisted of mixing graphite and the respective metal hydride powders in stoichio-metric proportions and subsequent heating of such mixtures in a hydrogen atmosphere in carbon crucibles. The heating was by high frequency to temperatures ranging between 1700" and 2100°C. The resulting carbide was then comminuted and screened to the desired particle size. ZrB, and TiB, powders were produced by the electrolysis of fused salt baths, according to the method described by Andriex.. The borides of niobium, vanadium, tantalum, molybdenum, chro-ium, and iron were obtained by mixing the respective metal and boron powders in the desired proportions. Such metal-boron mixtures were heated in a high frequency furnace to form boride powders. For each metal-carbon-boron group (Tables I1 through XI) a metal, its hydride, carbide or boride were mixed with carbon, boron or boron carbide powders. The additions of carbon, boron or boron carbide powders to any of these metals or metal compounds were calculated to satisfy a particular carbide or boride phase that according to the literature (Table I) had definitely been established by X-ray diffraction work. Samples of powder mixtures were hot pressed in graphite molds that were heated by direct conduction. The specimen dimensions were approximately 2.5X1X1 cm. Hot pressing temperatures were measured optically and maintained for approximately 30 sec under a constant pressure of about 1.3 ton per sq in. Wherever possible, an attempt to obtain maximum specimen density was made by temperature variation. Electrical resistivity testing was done by measuring potentiometrically the voltage drop over a length of 1.5 cm for a current of 10 amp, at room temperature. To obtain electrical resistivities for specific carbide or boride phases, values were plotted as a function of the respective sample densities

Jan 1, 1953

By Michel Dupuy, Daniel Calais

The study of self-diffusion of plutonium in E phase has been carried out by the welded couples method. The tracer used was puZ4O which is detected by its X-ray emission (conversion lines of uranium which are computed between 13 and 21 kev). Intensities were measured with a scintillation counter. Each layer was removed in a direction parallel to the original interface with a grinding machine and a thickness measured with a pneumatic comparator. The concentration-penetration curves obtained were corrected for the effect of heating time from room temperature to annealing temperature and for the expansion due to phase transformations of plutonium. They were analyzed by the generalized Gruzin method. Self-diffusion of plutonium in E Phase is very fast cm per sec between 500" and 620°C) and the diffusion zones are 2 to 3 mm wide for annealing times ranging from 30 min up to 10 hr. The Arrhenius law gives the temperature dependence in the form: From the point of view of self-dqfusion, PUE phase falls into the anomalous bcc metals category (Tip , Hfp, Zrp, Uy) with a low-frequency factor and an activation energy lower than those provided by standard correlations. No theory proposed hitherto to explain these anomalies (influence of dislocations, of extrinsic vacancies bonded to inlpurities, of bi-vacancies) can clearly explain the self-diffusion coeffzcients of plutonium. DIFFUSION in bcc metals is a present-day problem. A recent symposium (Gatlinburg, 1964), followed by a book,&apos; has been devoted to it. A great many experiments seem to show that diffusion in certain bcc metals obeys unexpected laws. The activation energies measured are sometimes strangely low (B hafnium, y uranium). For certain metals (0 zirconium, p titanium) the curves of log D (D = diffusion coefficient) as a function of 1/T (T = absolute temperature) are not linear. The frequency factors Do, which are of the order of 1 sq cm sec-&apos; in fcc metals, vary from 1 to 10~6 sq cm sec-&apos;. Various theories have been put forward to explain these anomalies; none is yet satisfactory. We wished to introduce a new experimental result by studying the self-diffusion in c plutonium. This allotropic phase, stable from 475°C up to the melting point (640°C), is in fact bcc. Unfortunately, nothing is known of the characteristics of the point defects in this phase, which limits the scope of the hypothesis which can be made about the mechanism(s) of self-diffusion in plutonium. 1) EXPERIMENTAL METHODS 1) Principle. We used the welded couple method. The two pellets of the couple initially have different 240 isotope contents (X emitter). After diffusion, the concentration/penetration curves are drawn up by the generalized Gruzin method. 2) Gamma Spectrography. The metal used in our study is plutonium, either low in puZ4O (isotopic content 1 pct) or high in puZ4O (8 pct). The latter also contains plutonium 241 (-1 pct) and 300 ppm of ameri-cium produced by the reaction Pu2U-AmM1 + 8-. The emission spectra of these two plutoniums placed in leak-tight vinyl bags have been studied by y spectrograph~. The detector is a thin crystal of thallium-doped sodium iodide. The activity of the plutonium rich in 240 is about twice that of the plutonium low in 240 in the energy band of 17 kev (L conversion lines of uranium); this band was used in these measurements. 3) Preparation and Examination of the Diffusion Couples. Diffusion couples were made from plutonium with a high and low PU"&apos; content. Pellets (6 6 mm. thickness 3 mm) mounted on a polishing disc with ground parallel faces were polished mechanically on both sides. In this way, pellets with two parallel faces were easily obtained. The polished pellets were joined by a 6 phase anneal (420°C, 1 hr) in a small screw press (pressure of 20 kg per sq mm cold); a centering ring enabled the two pellets to be pressed coaxially. The couples then were subjected to the diffusion treatment by annealing in the E phase in sealed silica ampules in argon at atmospheric pressure. The annealing temperatures and times are given in Table I. The couples were encased in a mild steel ring, the joint interface being thus parallel to the ground face of the ring. The diffusion couple/ring assembly underwent successive abrasions by means of a magnetic plate grinder. The thickness of the abraded layer was measured with a Solex pneumatic comparator when it was less than 0.1 mm (accuracy 0.2 p) or with a mechanical micrometer (accuracy 3 p) for passes of the order of 0.2 mm. All these operations were done in glove boxes, as plutonium is particularly toxic. After each abrasion we determined the emission spectrum of the ground face. The emissive surface is defined by means of a diaphragm 3 mm in diam. We noted more particularly the X activity in the 17-kev

Jan 1, 1969

By Arthur Baker III

The orebodies are long bedding-plane lenses of chalcopyrite and sphalerite, associated with garnetite masses. Most of the orebodies are within a 50-ft thickness of Cambrian limestone; other Paleozoic limestones and dolomites are locally metamorphosed but only slightly mineralized. Pre-mineral faults are numerous, but shallow folds were the main ore-localizing structures. JOHNSON camp is in the northwestern part of Cochise County, Ariz., about 50 miles east of Tucson. The nearest major mining districts are Tombstone and Bisbee, respectively 27 and 50 miles to the south, and the Superior-Miami-Globe-Ray porphyry copper group, about 90 miles northwest. Like many mining camps of the Southwest, Johnson camp is said to have been worked by the Spaniards. The first production on record was in the early 1880&apos;s, when an unknown amount of oxidized copper-silver ore in the Peabody mine was mined from replacement orebodies in the Pennsyl-vanian Naco formation. From 1904 to 1911 outcropping oxidized ores in the Cambrian Abrigo formation were worked, and an estimated 100,000 tons of copper ore were shipped. In 1912 the first large sulphide orebody of the district, the Republic Manto orebody, was discovered, and in the following few years some 250,000 tons of predominantly sulphide ore were shipped. The average grade of this ore was approximately 4.5 pct Cu, 6 pct Zn, 0.8 oz. Ag, and 0.001 oz Au. The Republic, Copper Chief, and Mammoth mines were the principal producers during this period. Mining ceased in the district after 1920, and until 1943 only small-scale leasing operations were carried on. In 1943 all the mines that had been productive were acquired by the Coronado Copper and Zinc Co., the present operators, and in the 10 years since that time the district production has amounted to approximately 350,000 tons of milling ore averaging 2 pct Cu and 6 pct Zn. Most of this ore was produced from the Republic and Mammoth mines, but since 1950 a large part of the production has been from the new Moore mine. The total known production from the district, then, is about 3/4 million tons of copper and copper-zinc ore of low grade. All of this ore was produced from orebodies associated with garnetite in the middle member of the Cambrian Abrigo formation. In addition to this known production, an unknown tonnage of ore was extracted from the Peabody mine orebodies that lie in the Pennsylvanian Naco formation. Published information on the geology of the district is limited. Aside from brief references in various mining journals, only three papers on the district have been published.1-3 One of these is a U. S. Bureau of Mines report on a diamond drilling program, one is a brief paper on the general geology of the district, and the third is a report on geochemical experiments, with a section on the occurrence of the orebodies. The last two are by John Cooper, of the United States Geological Survey, who has done much detailed work in the area. Stratigraphy The rocks of the mineralized area are Paleozoic sediments ranging in age from Cambrian to Penn-sylvanian. Several disconformities are present in the stratigraphic column, the most important one being between the Cambrian Abrigo formation and the Devonian Martin formation. There are no angular unconformities. Within the district, the Paleozoic sediments lie in a fairly uniform monocline, striking northwest and dipping 30" to 50" northeast. This local monocline is part of a domal struc-ture centered in the Little Dragoon mountains to the southwest. The Texas Canyon stock, a quartz monzonite body intruded probably during the Laramide revolution, lies south of the mineralized area (Ref. 2, p. 33). The Paleozoic rocks dip away from the stock, and on the surface are separated from it by at least 1500 ft of Pre-Cambrian rocks. The outcrop pattern of the northeastern edge of the stock suggests that it may dip gently northeastward, passing below the mineralized area at moderate depth. No quartz monzonite has been found in mine workings or diamond drill holes, which reach to depths of 1000 ft. The only igneous rock found in the mineralized area is a lamprophyre dike cutting the Naco limestone in and near workings of the Peabody mine. With the exception of the lowermost beds—the Bolsa quartzite and the shaly lower member of the Abrigo formation—the Paleozoic sediments are predominantly carbonate rocks, Fig. 1. The middle member of the Abrigo formation, which contains the principal ore-bearing beds, is limestone, with thin shale partings throughout most of its 250-ft thick-ness. Near the top of this member is a sandy bed some 25 ft thick. The upper member of the Abrigo formation and the lower half of the overlying Devonian Martin formation are dolomitic, with num-

Jan 1, 1954

By David C. Russell

The Department of the Interior used to be a quiet, noncontroversial, almost boring agency. It, after all is the fifth oldest of the Departments, and as an old line Federal agency it has studiously performed its preservation and resource management functions in a caretaker mode--though some would say more "undertaker" than "caretaker"--locking up the body and soul of America piece-by-piece. Yes, quiet, serene. That is until Jim Watt showed up. And we have all seen that version of Mt. Vesuvius which resulted--only it was the environmentalists who blew their tops. Ronald Reagan chose Jim Watt as Secretary of this fine old agency to prove that one-third of our Nation&apos;s land and over a billion acres on the Outer Continental Shelf can work for this Nation. At the foundation of President Reagan&apos;s charge to Secretary Watt is a belief in the tenets of the free enterprise system, and in the individual freedoms upon which this country was founded. There are those who don&apos;t share this belief in democracy and free enterprise, and those who believe this 205 year experiment called the United States of America will fail. Nikita Krushchev said "we will bury you"--obviously he didn&apos;t agree with our system. An Italian sociologist, Franco Ferrorotti, said bureaucratic stagnation will kill capitalism. Certainly we have all felt the ravages of bloated bureaucracies. Perhaps one indicator in the United States is the Federal Register, that daily compilation of Government&apos;s largesse. In 1970, 20,000 pages of the Federal Register were published. A decade later, in 1980, that volume had quadrupled to 80,000 pages. The Federal bureaucracy can stagnate from excessive budgets as well. The Interior Department spent $60 million on administering Federal coal leasing in 1981. That&apos;s nearly two bits a ton for every ton of coal leased in 1981. You wouldn&apos;t stay in business very long if your administrative overhead on inventory was that outrageous. But the pessimism of our critics is apparent from more than red tape and bloated budgets. For decades America has been fasting--consuming too little of America&apos;s wealth of minerals, subsisting instead on a diet heavily reliant upon mid-east oil, with little emphasis or concern for inventorying and developing domestic energy and mineral resources. Economics--yes. But short-term, short-sighted economics. Excessively dependent upon foreign imports, of oil, cobalt, chrome and other strategic minerals, the U.S. measures its time before another embargo--or fallen Shah, or Soviet manipulation, or Saudi shift, or, as we witnessed in Egypt, assassination--an untimely loss to mankind and efforts to bring peace to the troubled mid-east. These disruptions, in addition to their tragic human tolls, impair the free world&apos;s security. Huge chunks of the United States have been locked away in dozens of single land use categories in the name of conservation, with only the foggiest idea of what resources might be denied the American people-and this at a time of unacceptable levels of energy and strategic mineral imports. More than half and perhaps two-thirds of all Government-owned lands are totally withdrawn from or severely restricted to development under the mining and leasing laws. We must continue to rid Government of the overly zealous restraints which have been keeping us from drawing upon that which can help restore our economy and national security. When we assumed responsibility, the United States was dependent on foreign sources for about 40 percent of its oil. In 1981, our oil import bill was approximately $83 billion--nearly 17 times what it was in 1972. Our reliance on foreign sources for essential minerals is even more disturbing. We must look to other countries--some unfriendly, some unstable--for 22 of 36 strategically critical minerals. Yet the energy resources on federal lands which are owned by the American people could meet our needs for centuries if properly managed. Eighty-five percent of the crude oil yet to be discovered in America is likely to come from public lands, as will 40 percent of the natural gas, 35 percent of the coal, 80 percent of the oil shale, nearly all of the tar sands, and substantial portions of uranium and geothermal energy. Our vast hardrock-mineral wealth includes untapped deposits of essential elements we now import, such as chromium, copper, platinum, and cobalt. The obvious question is, if these abundant resources can help to revitalize our economic strength and to preserve our national security, why aren&apos;t we using them to better advantage? To a large extent, the answer can be found in past decisions to restrict public access to the federal estate, thus deferring to us or our successors the tough decisions that flow from Congress&apos; mandate to provide for environmentally responsible development of America&apos;s energy and mineral treasures. Here is the legacy this Administration inherited: In January 1981, 7 years after the onset of the Mideast oil embargo: ---Less than 15 percent of federal onshore lands were under lease for oil and gas development; ---No oil and gas leases had been issued in Alaska for 15 years;

Jan 1, 1982

By R. E. Powers, E. H. Kinelski, H. A. Morrissey

A group of 17 American blast-furnace sinters, an American open-hearth sinter, an American iron ore, and a Swedish sinter were used to evaluate testing methods adapted to appraise sinter properties. Statistical calculations were performed on the data to determine correlation coefficients for several sets of sinter properties. Properties of strength and dusting were related to total porosity, slag ratio, and total slag. Reducibility was related to the degree of oxidation of the sinters. THIS report to the American iron and steel industry marks the completion of a 1949 survey of blast-furnace sinter practice sponsored by the Subcommittee on Agglomeration of Fines of the American Iron & Steel Institute. The use of sinter in blast furnaces, sinter properties, raw materials, and sinter plant operation have been reported recently.1,2 After preliminary research and study," test procedures were adapted to appraise the physical and chemical properties of sinter to determine what constitutes a good sinter. During the 1949 to 1950 plant survey each plant submitted a 400-lb grab sample to research personnel at Mellon Institute, Pittsburgh, Pa. A 400-lb sample was also submitted from Sweden. In addition, 2 tons of group 3 fines iron ore were obtained from a Pittsburgh steel plant. The following tests were performed on the iron ore sample and on the 19 sinter samples: chemical analysis; impact test for strength and dusting; reducibility test; surface area measurements, B.E.T. nitrogen adsorption method; S.K. porosity test; Davis tube magnetic analysis; X-ray diffraction analysis for magnetite and hematite; and microstructure. Results of these evaluations are discussed in this paper and supply a critical look at testing procedures used to determine sinter quality. Sinter Tests and Results Each 400-lb grab sample of sinter was secured at a time when it was believed to represent normal production practice at each plant. It was not possible to use the same sampling procedures throughout the survey; consequently samples were taken from blast-furnace bins, cooling tables, and railroad cars. These were very useful for evaluation of test methods, since they were obtained from plants with widely divergent operations. With the exception of Swedish sinter and sinter sample N, which were produced on the Greenawalt type of pans, all survey sinters were produced on the Dwight-Lloyd type of sintering machines. Sinters submitted for test were prepared in identical manner by crushing in a roll crusher (set at 1 in.), mixing, and quartering. To secure specific size fractions for tests, one quarter of the sample was crushed in a jaw crusher and hammer mill to obtain a —10 mesh size. The remainder was screened to obtain specific size fractions. The group 3 fines iron ore was dried and screened and samples were taken from selected screen sizes to be used for various tests. Prior to testing, each ore sample except the —100 mesh fraction was washed with water to remove all fine material and was then dried. This iron ore, a hematitic ore from the Lake Superior region, was used as a base line for comparing results of tests on sinters. The iron ore did not lend itself to impact testing, since it was compacted rather than crushed in the test, and no impact tests are reported. However, the iron ore was subjected to all remaining physical tests to be described. Chemical Analysis: Table I presents chemical analyses performed on the survey sinter samples. Included in this table are data obtained from determination of FeO and the slag relationships: CaO + MgO and total slag (CaO + MgO + SiO, SiO2 + Al2o3 + TiO2). The percentage of FeO was used as an indication of the percentage of magnetite in the sinter. It was believed that slag relationships could be correlated with sinter properties. During initial determination of FeO great disagreement arose among various laboratories, both as to the results and the methods of determining values. Table I lists the values of FeO resulting from the U. S. Steel Corp. method of chemical analysis,&apos; which reports the total FeO soluble in hydrochloric and hydrofluoric acids (metallic iron not removed) with dry ice used to produce the protective atmosphere during digestion. Use of dry ice was a modification required to obtain reproducible results. In this method, the iron silicates and metallic iron are believed to go into solution and are therefore reported as FeO. This is important, for in the study of the microstructure of sinters, glassy constituents suspected of containing FeO as well as crystallized phases of undetermined identity which may also contain FeO have been observed. Strength Test by Impact: In evaluating sinter quality, one of the properties stressed most by blastfurnace operators is strength. This strength may be described as the resistance to breakage during handling of sinter between the sinter plant and the blast-furnace bins. It is also the strength necessary to withstand the burden in the blast-furnace. After

Jan 1, 1955

By W. H. Porter, M. E. Wadsworth, J. R. Lewis, K. L. Leiter

The oxidation of Cu2S in oxygen and the sulfating of Cu2O in oxygen-sulfur dioxide atmospheres was carried out under a variety of conditions. The oxidation of Cu2S was found to be retarded by entrapment of SO, and O2, which stabilized internal sulfates for long periods of time. The course of the reaction was followed by measuring weight changes and also by SO, evolution. Sulfating of Cu2O was a maximum at ratios of SO, to O2 approximating maximum SO, production. At elevated temperatures SO, was found to increase the rate of oxidation of Cu2 O to CuO even though sulfates did not form. All sulfating reactions followed the parabolic rate law indicating diffusion. MANY studies of the roasting of copper sulfides have been reported in recent years. Diev et al.1 investigated the roasting of chalcocite (Cu2S) in air, and oxygen enriched air. Lewis et al.2 also studied the oxidation of natural and synthetic chalcocite in air and oxygen atmospheres and their studies indicated that the maximum formation of water soluble sulfates occurred at approximately 450oC. Ashcroft3 reported that oxide production during the roasting of chalcocite resulted only from secondary decomposition of sulfates which were formed as primary products. peretti4 refuted this claim by showing that a layer of Cu2O appeared directly adjacent to the Cu2S during roasting of cylindrical briquettes of cupric sulfide, CuS. The linear advance of the Cu2S-Cu2O interface was used as a measure of the kinetics of the roasting reaction. The reactions proposed were: 2 CuS—Cu2S + 1/2 S2 [4-1] 1/2 S2 + O24 SO2 [4-2] cuzs +3/2 O2 4 Cu2O + SO2 [4-3] cu2o + 1/2 O2—2 cuo [4-4] At temperatures above 663oC, CuO was the only final solid phase reported. Below 663" C increasing amounts of sulfate were found mixed with the CuO. McCabe and Morgan5 investigated the roasting of discs of synthetic chalcocite and reported the following sequence of products beginning at the sulfide surface: Cu2O, a mixture of Cu2O and CuSO4, Cum,, CuO . CuSO4, and CuO. The principal reactions were reported to be: Cu2S + 3/2 O2-Cu2O + SO2 [5-1] CU2O + 2 SO2 + 3/2 O2—2 CUSO4 [5-21 2 CUSO4— CUO . cum, + SO3 [ 5-31 cuo . cuso4—-2 cuo + SO, [ 5-41 Eq.15-11 supports the claim of Peretti, Eq. [4-31, that CuzO is formed directly from Cu2S rather than as a secondary product from a sulfate as suggested by Ashcroft. On the other hand CuO was found to form as a secondary product from the decomposition of copper sulfate and basic copper sulfate, Eqs. [5-31 and [5-41. The formation of sulfates was explained by McCabe and morgan5 to be a direct reaction of Cu2O with 0, and SO, or SO, at distinct regions in which the partial pressures of each were such as to form the sulfate. Thornhill and pidgeon6 roasted both natural and synthetic chalcocite grains in air at temperatures between 420" and 550° C. They found a dense primary oxidation layer in contact with the sulfide. A secondary layer of porous oxidation products was found to expand with roasting time. The oxide products were leached away and the remaining core was studied by X-ray diffraction. The X-ray patterns showed an increased conversion of chalcocite to digenite with time. Digenite,7 a defect structure of cuprous sulfide, occurs naturally as Cu,-,S where x = 0.12 to 0.45, with an average analysis of Cu, ,S. The mechanism of digenite formation was proposed as: Cu2S + oxygen—Cu1-8S + 0.1 Cu2O [6-1] Cuj.eS + oxygen—0.9 Cu2S + SO2 [6-2] It is apparent from the above studies that the oxidation of Cu2S, ultimately ending in CuO, may be divided into ihree general stages (all of which may occur simultaneously): 1) primary oxidation to Cu2O; 2) secondary sulfate formation; and 3) sulfate decomposition. Consequently reactions of O2 and SO, with Cu20 constitute important aspects of the roasting of chalcocite. Virtually no studies have been made regarding sulfating reactions involving Cu,O. Mills and Evans8 noted the effect of sulfur dioxide on the oxidation of copper at low temperatures and low SO, partial pressures. They reported a measurable increase in the oxidation rate of copper when SO2, was present. Interest in the Cu2O-CuO-0, system has been limited predominantly to misciblllty studies and determinations of heats of formation by

Jan 1, 1961

By John Henderson

FOR some time now the most commonly accepted description of liquid silicate structure has been the "discrete ion" theory, proposed originally by Bockris and owe.&apos; This theory is that when certain metal oxides and silica are melted together, the continuous three dimensional silica lattice is broken down into large anionic groups, such as sheets, chains, and rings, to form a liquid containing these complex anions and simple cations. Each composition is characterized by "an equilibrium mixture of two or more of the discrete ions",&apos; and increasing metal oxide content causes a decrease in ion size. The implication is, and this implication has received tacit approval from subsequent workers, that these anions are rigid structures and that once formed they are quite stable. The discrete ion theory has been found to fit the results of the great majority of structural studies, but in a few areas it is not entirely satisfactory. For example it does not explain clearly the effect of temperature on melt structure,3 nor does it allow for free oxygen ions over wide composition ranges, the occurrence of which has been postulated to explain sulfur4 and water5 solubility in liquid silicates. In lime-silica-alumina melts the discrete ion theory is even less satisfactory, and in particular the apparent difference in the mechanism of transport of calcium in electrical conduction8 and self-diffusion,&apos; and the mechanism of the self-diffusion of oxygen8 are very difficult to explain on this basis. By looking at melt structure in a slightly different way, however, a model emerges that does not pose these problems. It has been suggested5" that at each composition in a liquid silicate, there is a distribution of anion sizes; thus the dominant anionic species might be Si3,O9 but as well as these anions the melt may contain say sis0:i anions. Decreasing silica content and increasing temperature are said9 to reduce the size of the dominant species. Taking this concept further, it is now suggested that these complexes are not the rigid, stable entities originally envisaged, but rather that they exist on a time-average basis. In this way large groups are continually decaying to smaller groups and small groups reforming to larger groups. The most complete transport data 8-10 available are for a melt containing 40 wt pct CaO, 40 wt pct SiO2, and 20 wt pct Al2O3. Recalculating this composition in terms of ion fractions and bearing in mind the relative sizes of the constituent ions, Table I, it seems reasonable to regard this liquid as almost close packed oxygens, containing the other ions interstitially, in which regions of local order exist. On this basis, all oxygen positions are equivalent and, since an oxygen is always adjacent to other oxygens, its diffusion occurs by successive small movements, in a cooperative manner, in accord with modern liquid theories." Silicon diffusion is much less favorable, firstly because there are fewer positions into which it can move and secondly, because it has the rather rigid restriction that it always tends to be co-ordinated with four oxygens. Silicon self-diffusion is therefore probably best regarded as being effected by the decay and reformation of anionic groups or, in other words, by the redistribution of regions of local order. Calcium self-diffusion should occur more readily than silicon, because its co-ordination requirements are not as stringent, but not as readily as oxygen, because there are fewer positions into which it can move. There is the further restriction that electrical neutrality must be maintained, hence calcium diffusion should be regarded as the process providing for electrical neutrality in the redistribution of regions of local order. That is, silicon and calcium self-diffusion occur, basically, by the same process. Aluminum self-diffusivity should be somewhere between calcium and silicon because, for reasons discussed elsewhere,&apos; part of the aluminum is equivalent to calcium and part equivalent to silicon. Consider now self-diffusion as a rate process. The simplest equation is: D = Do exp (-E/RT) [I] This equation can be restated in much more explicit forms but neither the accuracy of the available data, nor the present state of knowledge of rate theory as applied to liquids justifies any degree of sophistication. Nevertheless the terms of Eq. [I] do have significance;12 Do is related, however loose this relationship may be, to the frequency with which reacting species are in favorable positions to diffuse, and E is an indication of the energy barrier that must be overcome to allow diffusion to proceed. For the 40 wt pct CaO, 40 wt pct SiO2, 20 wt pct Al2O3, melt, the apparent activation energies for self-diffusion of calcium, silicon, and aluminum are not significantly different from 70 kcal per mole of diffusate,&apos; in agreement with the postulate that these elements diffuse by the same process. For oxygen self-diffusion E is about 85 kcal per mole,&apos; again in agreement with the idea that oxygen is transported,

Jan 1, 1963