Search Documents

By Harry A. Lipsitt, Douglas Y. Wang

A fatigue study in completely reversed axial tension-compression has been perforried on high-purity titanium and on three high-purity alloys of titanium. The alloys each contain approxi7nately 0.75 at. pct of a single interstitial element; carbon, nitrogen, and oxygen, respectivley. The results corroborate a previously published theory which proposed that strain aging under alternating stress was responsible for the fatigue limit behavior of certain alloys. The present data indicate that in these alloys an increasing strain-talline aging effect under alternating stress is provided by oxygen, carbon, and nitrogen, respectively. CURRENT research on the nature of the fatigue limit in metals suggests that the presence of a fatigue limit in metallic materials is a manifestation of strain aging that occurs under alternating stress.lm5 A comprehensive theoretical model based on the above hypothesis has been developed to explain the existence of a fatigue limit.' This model also provides increased insight into several other fatigue phenomena as under stressing, overstressing, and coaxing effects. The theory, as well, provides equal understanding for those cases where no real fatigue limit is observed. Briefly, this theory assumes that the S-N curve for a pure metal is a smooth function of the applied stress, and the effect of adding an element that is soluble (or forms a precipitate) in the pure metal is simply to shift the S-N curve to the right. If the added element confers the power to strain age, the result is a further shift of the S-N curve, this time upward and to the right. Since strain aging is not expected to be a strong function of stress, and since damage per cycle is known to be quite stress dependent, it is to be expected that there will be some limiting lower stress at which the strengthening due to strain aging will balance the damage due to crack propagation. This stress is the fatigue limit. The position of the fatigue-limit knee was thought to be a function of the magnitude of the strain-aging effect on both the finite and infinite life portions of the S-N curve. Although the strain aging hypothesis seems to be reasonably valid for bcc materials,2'6 it needed to be tested for both fcc and cph metals. This report is the first of a series concerning the fatigue-limit behavior of titanium with varying amounts of the interstitial solutes (C, N,, and 4) that are known to cause static strain aging in titanium. Yield-point effects have been reported for polycrys-talline high-purity titanium alloys containing either carbon, nitrogen, or oxygen.7'9 These effects were observed at testing temperatures in the range 100 to 300'. In addition yield-point and strain-aging effects have been reported for single crystals of titanium containing 0.1 wt pct C plus N.' These yield points were observed over a wide temperature range, but no room-temperature aging occurred. Aging at 180' was required to cause the return of the yield point. The magnitude of the yield phenomena in titanium containing interstitials is not expected to be as large as is observed in bcc metals because of several factors. Titanium has a very high chemical affinity for oxygen and nitrogen. The thermodynamic stability of solutions of oxygen or nitrogen in titanium is recognized. Lattice parameter measurements of titanium containing arbon, oxygen,1° or nitrogen" show that the "c" parameter is expanded more than the "a" parameter, but that up to about 2 wt pct this results in an insignificant change of the axial ratio 'c/a." Ehrlich" has shown that the sites occupied by interstitial atoms in titanium are spherically symmetrical and therefore a lattice expansion, at a constant c/a ratio, results in a simple dilation of the interstitial site. Such a dilation involving no shear has been shown to react only with edge components of dislocations.13 This causes only a weak pinning action. Shear stresses would be anticipated locally when only one of the two interstitial positions was occupied. The carbon atom will cause a symmetrical distortion of the lattice whereas the oxygen and nitrogen atoms have, in addition, the previously mentioned chemical affinity of titanium for these elements. These factors will result in a considerably smaller reduction of free energy upon the association of interstitial atoms with dislocations, and therefore a much weaker pinning than has been observed for the bcc metals. These considerations would lead to the hypothesis that of the interstitial elements considered here carbon would cause the strongest pinning effect in titanium where the amount of interstitial in solution is constant. This hypothesis will be borne out in the analysis of the present results.

Jan 1, 1962

By M. G. Jones

AT the Edwards mine a 4 by 81/2-ft two-compartment shaft is used for men and supplies in the older workings down to the 900-ft level. All ore is hoisted through a 5 by 15-ft three-compartment vertical shaft which extends to the 1500-ft level. Ore pockets are on the 1300- and 1500-ft levels. The two-compartment countershaft used below the latter level is 6 by 12 ft inclined at 42°. Ore from the countershaft is hoisted to the 1350-ft level where it flows by gravity to the 1500-ft level ore pocket. One-ton skips are used in balance in the countershaft and two-ton skips, also in balance, from the 1500-It level to the surface. These skips are slung under the man cages. Each level has an ore pocket of 200 tons' capacity. Development at Edwards is similar to that at Balmat except that haulage drifts are only 7 by 7 ft. They are mucked with Eimco-Finlay No. 12 loaders. A drill carriage similar to that used at Balmat is being tried out on the new 2300-ft level, the present bottom of the mine. Drifts are driven below grade and later raised 6 in. and ballasted with waste leaving a ditch 6 by 12 in. at one side for drainage. All development is done on an incentive bonus system, the miner paying for his explosives.

Jan 1, 1947

By Fred D. Fox

INTRODUCTION Approximately two years have elapsed since the first summary of the Grizzly Gulch Tailings Disposal Project was presented (1). Since that time, various physical modifications and additions to the existing treatment system have transpired. Further, Homestake and the entire mining industry have been impacted severely by additional modifications and additions concerning regulatory reform aimed at total disciplinary permitting and hazardous waste management. The following paper summarizes the past accomplishments relating to the Grizzly Gulch Tailings Disposal Project, correlating these accomplishments with future plans that are being developed to comply with regulations and to continue Homestake's commitment to environmental protection. The famous Homestake Gold Mine, located in Lead, South Dakota, has been operating in the Black Hills region for over a century, approximately 12 years before the south half of Dakota Territory was admitted to the Union as the State of South Dakota (2). During the 104 years of operational development, Homestake's Gold Mine, as well as the entire mining industry, have been influenced by numerous governmental regulations dating back to the Refuse Act of 1899 which regulated the deposit of debris from mines and stamp works. The enactment of the National Environmental Policy Act of 1969 may be thought of as the beginning of a long line of unplanned steps to deal with the many concerns of matters pertaining to the environment. Environmental awareness continued into the 1970’s, in which the Clean Air Act, Federal Water Pollution Control Act, Safe Drinking Water Act, Toxic Substances Control Act, and the Resource Conservation and Recovery Act (RCRA), to name a few, surfaced.

Jan 1, 1983

By Douglas R. Cook

The discovery of the gold deposits of the Jerritt Canyon district, with proven recoverable reserves in excess of 67.6 Mg (2.4 million oz) of gold, has contributed greatly to the tremendous resurgence in exploration for disseminated gold deposits in Nevada and throughout the west. The Jerritt Canyon project, including the Bell mine, is a 70130 joint venture between Freeport Gold Co. and FMC Gold Corp., both of which are wholly owned affiliates of the respective parent companies, Freeport-McMoRan Inc. and FMC Corp. The Jerritt Canyon district is located in the center of the Independence Mountains in north-central Elko County, Nevada, approximately 48 km (30 miles) northeast of the Carlin gold mine and 68 km (42 miles) north of Elko. Although the discovery of a commercial gold mine at Jerritt resulted from Freeport's exploration activities, the initial identification of a gold prospect in this area was a consequence of a program by FMC in the search for antimony. Their program was, in part, based on data in the Nevada Bureau of Mines Bulletin 61, "Antimony Deposits in Nevada." Work on the property began in 197 1 with geologic mapping followed by sampling and geochemical analyses. FMC geologists, including Robert Hawkins, Russell Hayden, and Hal Hurst recognized the striking similarity of the area to the geologic environment of Newmont's nearby Carlin gold deposit and soon focused on the gold potential. Increases in the price of gold served to stimulate interest and support for the exploration program. Geochemical samples were collected and analyzed for gold and trace elements usually associated with gold. A strong gold anomaly was found on the North Fork of Jerritt Canyon, which was called the Alchem anomaly. Drill-testing of this anomaly in 1973 revealed significant grades and thicknesses of gold mineralization in the lower portion of the Roberts Mountain formation. This initial discovery and subsequent close-spaced drilling proved the existence of several small pods of low-grade, gold-bearing material which cropped out at their up-dip edges, causing the surface anomaly. The mineralization was very encouraging, but was not, by itself, of economic importance from the standpoint of either grade or tonnage. Freeport Exploration Co. established a district office in Reno in late 1974 with responsibilities for exploration of hard minerals in the Basin and Range Province. Emphasis of the program was directed to the discovery of precious metal deposits. Enfield Bell, as District Manager for this new office, was particularly interested in the Independence Mountains and when the opportunity for a joint venture with FMC became available in 1976, it was aggressively pursued. Enfield Bell and David Stevens evaluated the FMC data and with the support of management formulated a bid in competition with 23 other companies. Freeport was selected by FMC as partner and a joint venture agreement was successfully negotiated. In mid-summer of 1976, Freeport Exploration Co. began an expanded program of detailed mapping and geochemical sampling under the direction of Bell and Stevens that led to new interpretations of the structural and alteration patterns. Hawkins' contribution during this period must be recognized since he not only was one of the original geologists assigned to the antimony and subsequently the gold investigations, but shortly after the start of the joint venture exploration, he was hired by Freeport Exploration Co. as the project geologist for the venture. Hawkins had studied the central Independence Mountains for his thesis and his understanding of the regional geology provided an excellent background for the continued exploration of the gold de- posits. His interpretation of the structure in the vicinity of the Alchem anomaly was of significant help in the Freeport exploration program. Drilling, based on this new under- standing of the geology, revealed the edge of the Marlboro Canyon ore body, which is now being exploited in the Bell mine. Although a classic bullseye geochemical target led to the discovery of the Alchem mineralization, the bulk of the reserves known today in the Marlboro Canyon area lay

Jan 1, 1985

By Winston W. Spencer

ALTHOUGH by far the largest A consumer of platinum metals in the world, the United States until recently has been in- significant as a producer. Writing in the "Minerals Yearbook" for 1939, H. W. Davis states that 1938 probably marked the beginning of an- other period of transition in the world production of the platinum metals. In that year, he says, their production in the United States jumped to 49,380 oz., or fourth in the rank of world output. In the ten-year period 1925 to 1934 United States output ,averaged only 8300 oz. Practically all of this sudden in- crease in production is attributable to the introduction of mechanized equipment at the placer deposits in the Goodnews Bay district of Alaska. This district adjoins the Bering Sea and Kuskokwim Bay (accented on the first syllable), and is approximately 120 miles south of the mouth of the Kuskokwim River. The nearest village and post office is Platinum, which lies about eleven miles north- east of the Goodnews Bay Mining Company's camp. A good gravel road connects the mine camp and the village.

Jan 1, 1940

By D. A. Augenstein, L. V. Amundson

Operators of coal preparation plants use equipment such as large-capacity dump trucks and bulldozers to haul, spread, and compact refuse material to conform to federal and state regulations governing the disposal of solid waste. For example, federal regulations require that refuse be spread and compacted in layers not more than 0.6 m (2 ft) thick. If the refuse layers or piles have poor load-bearing characteristics, movement of equipment over them becomes extremely difficult. The increasing cost and limited space for the construction of settling ponds extensively used for fine refuse disposal are leading to the growing practice of dewatering the fine refuse by means of vacuum disk filters and combining the resulting filter cake with coarse refuse for waste bank disposal. However, this filter cake usually contains more than 25% moisture. Therefore, when it is combined with coarse refuse, the moisture level of the total refuse may become high enough to impair the load-bearing strength of the refuse. Precisely this problem and the attendant difficulties of moving heavy equipment arose when vacuum disk filters were installed at a preparation plant in West Virginia. Combining the moist filter cake with the coarse refuse generally results in a 12% moisture level for the total refuse at this plant. At this moisture level, the refuse can still be handled with some dif¬ficulty by the equipment. However, the problem has frequently been compounded after rains. The experimental program described in this paper tested the following methods of improving the load-bearing properties of coal refuse: moisture reduction, addition of crushed coarse refuse, addition of fly ash, addition of lime, and addition of a mixture of lime and fly ash. In terms of a balance between economic and technical considerations, the most effective method was demonstrated by laboratory and plant tests to be a 2% to 5% addition of lime. Test Program During formulation of a test program aimed at improving coal refuse stability, a number of treatment techniques were selected for evaluation of their effect on refuse load-bearing characteristics. The reduction of refuse moisture was the first technique considered, since moisture content plays a key role in the bearing strength of a bulk solid. The reasoning was that a very small reduction in refuse moisture, provided it could be accomplished with minimal effort and cost, would be sufficient for eliminating refuse disposal problems most of the time. At the same time, it was recognized that periods of rainy weather would offset moisture-reduction measures. Other techniques selected for evaluation because of their potential for improving refuse-bearing strength at reasonable cost were: addition of crushed coarse refuse, the addition of fly ash, and the addition of lime. A simple one-dimensional laboratory compaction test was designed for evaluating the effect of each of these techniques on refuse-bearing strength. Those techniques that gave best results in the laboratory would then be tested in the preparation plant in West Virginia. Laboratory Tests The laboratory test for evaluating each of the techniques involved the application of weight to a load module placed on a refuse sample in a container. The distance the load module sank with an increasing amount of applied weight was measured, and the relationship between module displacement and loading was obtained. The refuse container was large enough to minimize the influence of wall effects, and to decrease the effect of the container bottom, a test was terminated when the load module sank to about three fourths the depth of the container. Initially, an attempt was made to conduct compaction experiments with samples of total plant refuse. However, the presence of + 0.13 m (+ 5 in.) material in the refuse interfered with the mechanics of the test, and the large amount of material required for a representative sample made testing extremely tedious and time-consuming. To avoid this problem and because it was believed that only the fine components of the refuse were the major cause of the stability problem, tests were conducted with 13 mm X 0 (1/2 in. X 0) material from the total plant refuse. The equipment consisted of a 0.46 m (18 in.) cubic container, a loading module with 62 cm2 loading area, weights totaling 450 kg, and a portable concrete mixer for combining water and additives with the refuse. Refuse samples of about 110 kg dry weight were used for each test. Experimental compaction test data were evaluated by plotting the linear displacement of the load module vs. loading. During compaction tests, the vertical displacement of the load module was measured at each loading value. The plot of

Jan 1, 1980

By Arthur B. Cummins

Diatomite is a siliceous rock of sedimentary origin, which may vary in degree of consolidation, but which consists mainly of the fossilized remains of diatoms. These are microscopic algae of the order Bacillaria. Diatomite is a unique form of silica, there being nothing else found in nature or produced by the manufacturing arts, that has its characteristic physical structure. There are numerous sediments which contain some diatom residues, admixed with a substantial amount of other materials-clay, carbonates or silica. While these are classified as diatomaceous silts, shales, or mudstones, they are not properly diatomite, a designation which is restricted to material of such qualities that it is suitable for commercial uses. The terms diatomaceous earth and kieselgur are synonymous with diatomite. The expressions infusorial earth and tripolite are obsolete. Diatomaceous silica is a definition best used to distinguish the basic component of diatomite, that is, the substance of the silica shell itself and is the major constituent of beneficiated diatomite or processed diatomaceous products. Some highly indurated materials, such as cherty shale, porcelanite, and opaline cherts which have resulted from diagenetic changes cannot be considered as diatomite although they may have been derived therefrom. On the other hand, diatomaceous oozes or loose sediments consisting mainly of diatom residues with or without much organic matter are source materials in process of becoming diatomite. Composition and Properties Pure diatomite is composed of opaline or hydrous silica, with small amounts of associated inorganic components-alumina, iron, alkaline earths, alkali metals, and minor constituents, all of which may occur in solid solution or as an essential part of the opal complex. Crude diatomite on the other hand may contain variable amounts of organic matter and soluble salts, and the usual contaminants of sedimentary formations-particles of rockforming minerals, sand, clay, or carbonates. As occurring in nature, diatomite contains an unusual amount of free water, which may vary from 10 to 60 pct or more. Diatomite, in the dry deposits which are characteristic of most of the important world deposits, occurs as a lightweight, comparatively light colored, "chalklike" material. The apparent density (dry basis) in situ varies from 20 to 40 lb per cu ft. The structure may be either massive or finely stratified. The earth is highly porous, usually soft and easily friable. It is nonreactive with acids and may be tentatively identified by the above characterizations. Confirmation is by microscopic examination which readily reveals the nature of the material. In deposits which contain organic matter the appearance of the material is totally different. The color may vary from light tan or gray to brown, greenish or nearly black. There is frequently difficulty in distinguishing it in the field from peatlike materials, but the difference is readily apparent after calcination. Powdered diatomite or diatomite products have an apparent density (loose weight) of 5 to 16 lb per cu ft, wet densities of 10 to 30 lb per cu ft, specific gravities 1.9 to 2.35, index of refraction 1.42 to 1.48. Diatomite is essentially micro-amorphous, but does show some tendency toward orientation in structure after a cristobalite pattern. Amorphous silica has been considered as a random association of cristobalite crystallites, the size of which are only slightly larger than a unit cell.20 The prop-

Jan 1, 1960

By A. D. LeClaire

This paper claims to provide a description of diffusion processes in solids which is said to be a reformulation of Fick's Laws and therefore, presumably, is supposed to be as general as those laws. In fact, it appears to provide no more than an atomic model of a rather restricted type leading even to results in conflict with experiment. The restriction lies in the assumption, carried throughout the paper, that even in a chemical gradient the jump frequency i of an atom of species i is a function only of the composition on the plane from which it jumps, i.e., ?i = ?i(cl, c2 ...). This means that atom jumps in opposite directions from a site would occur with equal probability. There would then be no tendency for atoms to migrate preferentially in one direction along the gradient, so the model is incapable of accommodating the well established phenomenon of atom drift in a chemical gradient.10-" A more general model that does not suffer this difficulty was discussed in a paper13 I wrote in 1958, and the results therein already contain as a special case those given by Dr. DeHoff. The simplest sufficiently general assumption is to suppose that ?i is a function ?i(a1 a2,...; wl, w2, ..-) of the concentration aj and ? ?j at the site from which and to which, respectively, a jump is made. Any such function can always be expressed as a function of the sum and difference of its arguments. Thus d is the jump distance and g a geometrical constant. Obviously this is zero, i.e., no drift, if ??ri/??j = 0. X is written in this form, revealing two components of the total drift, because it can be shown that only the second component, depending on ??'i/???(?j -aj), actually contributes to the net diffusion flux Ji. This component can be shown,13 by a straightforward ther-modynamic transcription, to be responsible for the so called "thermodynamic factor" in the Darken approximation for a chemical diffusion coefficient is the activity coefficient of i. Because the jump rate ?i in DeHoff's approximation has no dependence on the concentration gradient it does not seem possible to derive the Darken equation from his Eq. [6] without the further rather arbitrary assumption that ?i = const ri. It is to be noted too that the author's derivation of Eq. [12] is only valid when ?i is in fact a function only of the cj at the plane from which the atom jumps—for only then is ci?i a unique function of position and therefore of A. Author's Reply R. T. DeHoff I am indebted to Dr. LeClaire for his penetrating discussion of my paper, "A Reformulation of Fick's First Law". I concede that it does appear initially that the relationship presented in my paper could be derived from the detail atomic model of random walk and atom drift presented in Dr. LeClaire's approach. However, there are some important differences in concept and viewpoint which deserve further amplification. Dr. LeClaire's description is penetrating, detailed, and fundamental. His approach describes an extremely complex phenomenon, the displacement of atoms in a multicomponent gradient, at a sophisticated level. For precisely these reasons, his analysis has not proved to be useful for experimental studies that have practical objectives. For example, in order to obtain the information required to complete the analysis of the jump frequency distribution in a ternary system, the following series of experiments would, I believe, be required: 1) chemical (gradient) experiments, covering the composition range;

Jan 1, 1970

By D. E. Light, J. Y. Somnay

The use of alkyl acid phosphates and their alkali salts as collectors for the uranium minerals brannerite and uranothorite was investigated. In particular a detailed flotation study was carried out using isooctyl acid phosphate as a promoter. Tests were conducted on an Elliot Lake ore analyzing 0.11% U 3 O 8, chiefly as brannerite, and approximately 9% pyrite. Pyrite was floated with conventional reagents; the flotation of brannerite from the sulfide tails was studied to determine the optimum pH and pulp density to employ with isooctyl phosphate. 4 uranium recovery of over 90% was possible with a selectivity index of 8.95 when these tails were floated at 17% solids and a pH of 1.7. The effect of various modifying agents was also investigated. Aluminum sulfate, lactic acid and sodium silicate improved the selectivity of collection yielding selectivity indices of 9.5 to 11.5. Ferric chloride was found to be a depressant for brannerite. A Bancroft ore which analyzed 0.13% U308, and contained uranothorite was also amenable to flotation using isooctyl acid phosphate. The composite uranium concentrate obtained assayed 0.41% U308 at a recovery of 95.3% and a ratio of concentration of 3.24. Previous investigations1,2 on the flotation of brannerite, pitchblende, uraninite and/or uranothorite from various Canadian ores showed that fatty acids, including oleic acid and tall oils, petroleum sulfo-nates and alkyl acid phosphates, were suitable collectors for these uranium minerals. Additional work carried out by the Mines Branch in Ottawa on brannerite ores from the Elliot Lake area3 used the tall oil Acintol FA-1, or FA-2, as a collector. A recovery of 92% of the uranium in a concentrate which contained about 55% of the weight and assayed 0.2% U3O8 was obtained from a feed of 0.1% U308 in this latter instance. Eigeles et a1.4 carried out flotation on pitchblende ores and found that the most promising results were obtained with alkyl acid phosphates and their alkali salts as collectors. Within the homologous series of alkyl phosphates, the best flotation characteristics were displayed by isooctyl phosphates. Further, isooctyl phosphate derivatives were much less sensitive to hardness forming salts than oleic acid, enabling all flotation tests to be made with tap water. This paper describes tests employing alkyl acid phosphates for the flotation of ores containing brannerite and uranothorite. The flotation work described herein was conducted in conjunction with the development of a leaching process for Elliot Lake, Ontario, brannerite ores. In this process pyrite in the ore was concentrated by flotation and roasted, providing sulfur dioxide which was used to produce sulfuric acid in situ in a liquid-solid extraction process. A 40% conversion to sulfuric acid was indicated at a leaching temperature of 80°C; this resulted in the extraction of 93% of the total uranium after a contact period of 6 hr. A rotary kiln was used as a leaching vessel. Preconcentration by flotation was studied in conjunction with the above process to reduce the bulk of solids for leaching and, consequently, capital expenditure. The materials and methods used in this study will be described initially. Then the results will be presented in two sections, the first dealing with the brannerite ore, and the second with the uranothorite. Finally an economic analysis of the results on the brannerite material will be made. MATERIALS AND METHODS The two ore samples examined were from uranium-producing mines. One was a brannerite ore, a uraniferous quartz pebble conglomerate from the Elliot Lake area, which contained 0.11% U3O8, 4.3% S and 4.5% total Fe. Mineralogical studies of similar samples indicated the presence of quartz, microcline, sericite, pyrite and minor quantities of rutile, pyrrho-tite and zircon. The other ore was from a biotite-rich pegmatite in the Bancroft area; feldspar, quartz and biotite formed the major gangue components. Uranothorite was the principal uranium mineral; the head assay was 0.13% U 3 O 8. Preparation of Feed: The ores were crushed to —10

Jan 1, 1963

The United States has never been a large coal exporting nation, and normally such shipments, aside from those to Canada, have been relatively small. Since the first railroad was built across the Allegheny Mountains, the imports have been quite small also, although fairly large tonnages of Canadian coal have moved into the extreme northeastern and northwestern states at times. In Table 21 are shown all data available about these shipments from 1790 to 1939, inclusive.

Jan 1, 1942

By Howard A. Meyerhoff

AS a source of mineral wealth, the larger islands of the West Indies have never had an enviable reputation. The Spaniards took possession of them in the sixteenth century hopeful that they would yield gold; but it was not long before they knew the rather rigid limitations of Antillean gold production and were seeking-and finding ---richer deposits of the precious metals in Mexico and on the west coast of South America. Although a great deal of wealth changed hands in the West Indian islands in subsequent years, the fact may be laid to the abundance of pirates and privateers and not to the richness of local ore bodies.

Jan 1, 1941

By V. GUDKOV

THE iron-ore deposits of Russia were estimated at 2,200,000,000 by the Russian Geological Survey, in 1910; but this estimate must be considered as being far too low. The estimate for Siberia, which had been explored very little for iron ore, was placed at about 27,000,000 tons. The exploration, in 1912-14, of the three iron deposits-Blagodat, Baakal and Kamensk-raised the estimate of the iron-ore reserves or the Urals from 282,000,000 tons to about 400,000,000 tons. Conservative estimates of the explored deposits made by the writer in 1915, gave the following figures: Southern Russia 900,000,000 tons, Ural Mountains 500,000,000 tons, Central Russia 900,000,- 000 tons, Siberia 100,000,000 tons, Caucasus 20,000,000 tons, or a total of 2,420,000,000 tons of iron ore, or 1,200,000,000 tons of pig iron.

Jan 1, 1921

By Carrol M. Beeson

Reasons are given for using a Boyle's-law porosimeter in conducting core analysis for either routine or research purposes. Among other things, it is pointed out that such a porosimeter permits the measurement of all basic properties on the same sample, thereby eliminating the sources of error inherent in the use of adjacent samples. References are made to investigations of gas adsorption on various porous materials, to show that the use of helium in Boyle's-law porosimeters reduces to negligible proportions the error due to the adsorption or desorption of the operating gas. Two Boyle's-law instruments are described. which permit accurate and rapid measurements of porosity. Schematic sketches and explanation:; are included, along with derivations of equations required in performing precise determinations. Summaries of data obtained during calibration are tabulated and analyses of the data are resented as indications of the precision and accuracy of each device. Comparisons are also shown for measurements made with each of the instruments on the same test pieces and cores. INTRODUCTION An accurate porosimeter, operating on the principle of Boyle's law. is of considerable value in the analysis of cores for either routine or research purposes. This is due primarily to the fact that the measurement of porosity with such an instrument leaves the sample free of contamination by any liquid. When used in conjunction with an extraction apparatus' for determining oil and water saturations, a Boyle's-law porosimeter permits the measurement of all basic properties on the same sample. This eliminates the sources of error inherent in the use of adjacent samples, or the necessity of determining porosity after all other properties have been obtained. Large errors may result from combining measurements made on adjacent samples in order to obtain a single property. This type of error is definitely involved when oil and water are measured with one sample, and the pore vo1ume is measured with an adjacent one. Furthermore, the source of error would be present to some extent, even if the analyst could choose the samples so they were truly adjacent from a geological standpoint. The use of adjacent samples in routine core analysis also necessarily decreases the probability of correlating core properties. For example, the chance of correlating the "irreducible" interstitial-water saturation with permeability, is bound to be greatly reduced by measuring these properties on "adjacent" samples. For research purposes, amplification is scarcely required concerning the greater flexibility of a method for measuring porosity which leaves the core free of contamination by any liquid. Even under those circumstances which require that the core be saturated with a liquid, a previous measurement of porosity with a gas is useful in determining the degree of saturation that has been attained in the process. Furthermore, for comparable accuracy, porosity usually may be determined more rapidly with a gas than with a liquid. This advantage of the Boyle's-law instrument is most outstanding when the determination time is compared with that required in obtaining porosity by evacuation of the core followed by saturation with a liquid of known density. Several porosimeters which operate on the principle of Boyle's law have been described2,3,4,5,6,7 in the literature. No comparison will be attempted between those instruments and the ones described herein. Before helium gas became readily available, Boyle's-law porosimeters were subject to an appreciable error due to the adsorption of the operating gas on the surface of the core solids. There is considerable direct and indirect evidence in the literature to support the contention that the adsorption of helium on porous solids is negligible at room temperature. In discussing the use of Boyle's-law porosimeters, Washburn and Bunting2 stated that "for most ceramic bodies dry air is a satisfactory gas, but hydrogen will be required in some instances. Helium could, of course, be employed for all types of porous materials at room temperatures or above." Howard and Hulett8 obtained evidence that the adsorption of helium was negligible at room temperatures, even on activated carbon ; for the density measured with this gas was unaffected by changes in pressure or by changes in temperature from 25 °C to 75 °C. For oil-well cores, Taliaferro, Johnson, and Dewees" obtained lower porosities with helium than with air, but apparently did not study helium adsorption. From the work of these investigators, it follows that the use of helium in Boyle's-law porosimeters reduces the error due to gas adsorption to negligible proportions. This makes it possible to construct instruments which permit the determination of porosity with (1) a high degree of accuracy, (2) with great rapidity, and (3) without contamination. THE KOBE POROSIMETER The fundamental design of the Kobe Porosimeter was developed by Kobe, Inc., which firm built about 12 of the instruments during 1936 and 1937. Since that time, seven or eight more have been constructed with their permission, making a total of about 20 that have been put into operation.

Jan 1, 1950

By B. K. Larkin, H. R. Bailey

One of the most complicated and potentially one of the most promising secondary recovery methods is that of underground combustion. A number of field tests1,2,3 have been performed recently, apparently indicating that the underground combustion process is technically feasible. Relatively little has been published concerning laboratory or theoretical investigations of the underground combustion process. A few results of laboratory tube runs have been published4,5 and these data are valuable in understanding the process. The tube run data also emphasize the complexity of the thermal recovery process. Under various conditions, problems in heat transfer, multiphase fluid flow and combustion rates arise. For example, at low temperatures the combustion process proceeds so slowly that the front cannot be propagated. It is a problem in heat transfer to predict what injection rate of air is necessary to sustain combustion in a particular reservoir. Certain fluid flow aspects of underground combustion have been studied by Wilson, et a13. In the development of some of the flow equations, assumptions of adiabatic combustion were made. Vogel and Krueger1 have devised an analog model of the heat conduction process for a constant temperature cylindrical source moving radially outward in a homogeneous conducting medium. Jenkins and Ramey7, in a discussion of the Vogel and Krueger paper, present an analytical solution of a heat conduction problem related to problems considered by Vogel and Krueger. In the present paper an idealized model of the heat conduction problem is considered, which is more general than the model of Vogel and Krueger or Jenkins and Ramey. In particular, initial well heating, vertical heat losses and arbitrary frontal velocities are included. Equations for the temperature are obtained in terms of integrals. These integrals have been evaluated numerically for the case of frontal movement corresponding to a constant air injection rate and the case of constant frontal velocity. These results are presented graphically. From these graphs one may estimate conditions under which nearly adiabatic temperatures are obtained. The application of these results requires some knowledge of reservoir thermal properties, combustion rates and residual fuel content. Assuming no vertical losses it is possible to obtain explicit evaluations of these integrals in the linear con- stant velocity case and in the radial case corresponding to a constant air injection rate. PRESENTATION OF RESULTS The geometry of the radial system is illustrated in Fig. 1 where the reservoir is represented as a slab of thickness h. By making an energy balance for an infinitesimal volume, one obtains the following equation. T= x T+F(z.r).........(1) where F (z,r) =HWv1/pc (r-r1), if z h/2 = 0, if z >h/2 . In Eq. 1, V- represents the Laplacian operator and 8 is the Dirac delta function. Eq. 1 is the well-known diff-sion equation with a source term and may be applied to a linear system. For example, if the source is moving in the direction of the x axis, then the argument of the delta function is x - 1,. In deriving Eq. 1 the following approximations were made concerning the actual combustion process: (1) all heat transfer is by conduction in a homogeneous medium, (2) the formations bounding the reservoir have the same thermal properties as the reservoir, (3) the combustion fuel is consumed instantaneously as the front arrives, (4) there is a constant fuel concentration

By F. P. Nichols, F. A. Renninger

At one time, mineral aggregates were considered physically and chemically inert substances whose junction was that of an inex-pensive, easily obtainable source of bulk. This philosophy applied regardless of the use to which the aggregate was put. Research approaches dealing only with such gross physical properties as hardness, toughness, impact resistance, etc., dominated the early growth years of the industry and influenced specifications to a large extent. The accelerated pace of highway construction, which began toward the end of the depression years, focused attention on the need for extensive aggregate research. The realization in the early 1940's that aggregates interacted chemically as well as physically with their environment provided the stimulus necessary to translate the drab, systematic aggregate technology of the early 20th century into the changing, dynamic, challenging field it is today. Current research problems are discussed in detail. These efforts being pursued by contract research organizations, universities, governmental agencies, and industry transcend many fields. The mineral aggregate industry is the world's largest mining and processing endeavor, yet it is graced with little of the romance or public awareness that accompanies mention of gold, silver, iron, copper, uranium, or aluminum mining efforts. Much of this reaction is undoubtedly due to the fact that the uses to which crushed stone and sand and gravel are normally put are not fully recognized or appreciated by the general public. Almost 2 billion tons of mineral aggregates are quarried or mined, processed, and sold yearly in the United States alone.= The largest uses to which this vast tonnage is subjected are as ingredients of portland-cement concrete, bituminous concrete, and dense-graded aggregate-base course mixtures in the highway and building construction fields, two industries closely related to the expenditure of public funds. It is for these uses, partially in justification for the expenditures of vast sums of public funds, that mineral aggregate research truly finds its emphasis. The research efforts expended relative to the highway and building construction uses to which mineral aggregates are subjected are those of primary concern to the industry. These efforts, while of direct consequence to the industry due to the volume market represented, are of interest as well to public agencies, universities, and contract research firms. All actively conduct research, exchange information, and make contributions to the overall knowledge relative to aggregate usage and limitations. The research efforts of the mineral aggregate industry is the subject to be discussed throughout this paper. It covers all aspects of highway and building design, construction, and technology. The effects of research pertaining to quarrying and mining methods, blasting and processing techniques, storage and handling facilities, and equipment employment and development have left their mark on the crushed stone and sand and gravel industries. Such specialized efforts have resulted in dramatically increased production rates which in turn have served to keep costs, and hence prices, down. Despite the many inflationary pressures over the years, the cost of a ton of aggregate today is little more than it was 25 or 30 years ago. Mention is made of these efforts simply to emphasize the diversity of research in the mineral aggregate field. The remainder of this paper focuses on the complexity of what might be termed "product knowledge" research.

Jan 1, 1971

By W. A. Backofen

THE texture of cold-drawn copper wire has been described most often as a composite of [Ill] and [loo] directions aligned parallel to the wire axis.' Hibbard2 suggested that sufficient reduction will result in a single [111] texture and reported that after a reduction in area of 96.4 pct, such a texture was nearly obtained. Other work3-4 has shown that recrystallization within the temperature range of 300" to 1000°C does not alter the texture established by cold-drawing. In general, however, the texture of cold-drawn end annealed copper wire is not as well defined as the deformation texture. In the work reported here it was found that the double-fiber, deformation texture persists even after a reduction in area of 97.3 pct, and after recrystallization the texture is practically all [l00]. The wire used in this work was prepared in the following way. A 3/4-in. diam length of annealed OFHC copper was cold-drawn to a diameter of 1/8 in., a reduction in area of 97.3 pct. Each step in the drawing operation brought about a reduction in area that varied from 6 to 10 pct. Short lengths of this wire were etched in a 50-pct solution of nitric acid to a diameter of approximately 0.01 in. The texture of the cold-drawn specimens was determined with the conventional X-ray transmission diffraction technique using filtered copper radiation. The X-ray beam was directed perpendicularly to the wire axis and the diffraction rings were recorded on a flat film. Fig. la is typical of the photo-grams obtained and reveals that the cold-drawing schedule described above introduces a distinct, double-fiber texture with both [Ill] and [l00] directions aligned parallel to the wire axis. Supplementary measurements with a Geiger-Mueller counter of the intensity of reflection from (111) planes showed that the [111] component was much more predominant than the [l00]. Additional specimens of the drawn and etched wire were annealed in a salt bath at temperatures from 300" to 450°C. The texture of the recrystal-lized wire was studied with the same technique applied to the cold-drawn wire. Fig. lb is an X-ray photogram of a specimen annealed for 1 ½ at 300°C and is typical of those obtained from re-crystallized specimens. The photogram shows that recrystallization of this material is accompanied by the almost complete elimination of the [111] component of the deformation texture. A pronounced strengthening of the [100] component was evident from intensity measurements with a Geiger-Mueller counter. In summary, the deformation texture of cold-drawn OFHC copper wire, after a reduction in area of 97.3 pct, may still be described as a composite of [Ill] and [loo] directions aligned parallel to the wire axis. Recrystallization of such wire at temperatures from 300" to 450°C introduces a texture which consists almost entirely of a [loo] alignment. Acknowledgment Sidney Rolle of the Scomet Engineering Co. generously supplied the OFHC copper used in this work. References 'C. S. Barrett: Structure of Metals. p. 382 (1943) New York. McGraw-Hill Book Co. 2 W. R. Hibbard, Jr.: Trans. AIME (1949) -185, p. 598; Journal of Metals (Sept. 1949) TN 19E. 3 P. 422 of ref. 1. 'E. Schmid and G. Wasserman: Ztsch. Physik (1926) 40, p. 451.

Jan 1, 1952

By R. P. Abendroth

This study is concerned with the kinetics of selective oxidation of chromium in a commercial Fe-Cr-Ni alloy. Selective oxidation of chromium in this alloy, by use of a low oxygen-potential atmosphere, leads to the formation of a compact, protective layer of Cr2O3. This layer serves to protect this alloy from gross scaling when it is subsequently exposed to severe oxidizing conditions at high temperature. The interaction of low oxygen-potential atmospheres close to equilibrium with Fe-Cr and Ni-Cr alloys has been studied by others."' These studies werk concerned with the surface-structure variations under slightly oxidizing conditions. NO detailed study was made of the oxide scale formation kinetics, however. The alloy samples were cut from 0.012-in.-thick sheet, with an apparent surface area of 7.1 sq cm. These sheet samples were abraded through 4/0 metallographic paper, and washed in alcohol and acetone. The analysis of the sheet alloy is (in weight percent): 42 pct Ni, 5.5 pct Cr, 0.09 pct C, 0.18 pct Al, 0.36 pct Si, 0.26 pct Mn, balance Fe. The weight gain vs time data were obtained with a 2-g capacity fused-silica spring—cathetometer system. The spring deflection was optically magnified ten times before being read by the cathetometer. A sensitivity of about 0.01 mg was attainable. The spring was enclosed in a water jacket maintained at 60°C to minimize the effect of temperature changes. The sample was suspended from the spring with a fused-silica hangdown and was positioned in the thermal center of a mullite furnace tube. Sample temperature was read with the aid of a thermocouple placed outside the mullite tube and an inside vs outside temperature calibration. Temperature change during the course of a run was ±0.25°C, with a temperature gradient of less than l.O°C over the length of the sample. Total temperature uncertainty was no more than ±3.0°C. Alignment difficulties between the hangdown and radiation shields in the furnace tube required that the sample be positioned in the furnace when cold, and heated with the furnace until the temperature stabilized at the desired point. This required 5 to 6 hr, and was carried out in Matheson ultrahigh-purity hydrogen. Oxidation was started, after evacuation, by introducing a hydrogen-water vapor mixture, obtained by saturating hydrogen with water vapor at 31.00o ± 0.02oC. Oxidation was continued for 90 min. Gas flow was 300 ml per min during heat up and oxidation. Since only several milligrams of oxide are formed on each sample, chemical analysis of the oxide is difficult. A representative analysis is: 80 pct Cr2O3, 5 pct Fe2O3, 3 pct Al2O3, 4 pct MnO, 7 pct SiO2, 1 pct or less NiO. X-ray diffraction analysis of the oxide as formed on the alloy gives rhombohedra1 Cr2O3, and barely distinguishable amounts of a cubic spinel phase, and possibly AlZOs and SO2. The identification of these latter two compounds is by no means certain. The spinel phase could be based on iron or manganese as these elements are present in significant amounts in the oxide. The results of the kinetic studies using 31°C dew-point hydrogen-water vapor mixtures were found to conform to a parabolic rate law. In many cases the parabolic plot consisted of two intersecting straight lines, defining an early and a late rate for a particular run, and in the other cases the parabolic plot consisted of one straight line for the entire run. The slopes of the various straight lines were determined by the method of least squares. Reproducibility of the data was good enough for multiple runs at the same temperature such that the value given for the rate constant is the average for two or more closely similar values, rather than widely varying values of the rate, where more than one determination is indicated in Table I. The exhibition of only one or of two rate constants during a run can happen at the same temperature. Thus, Table I shows that at 11'74°C a single rate constant was obtained from one sample, while other samples oxidized at the same temperature gave an early and a late rate constant. It should be noted that the single rate constant corresponds very closely with the early rate constant. This is also true at 1153°C. The time at which the late rate started to appear was variable, usually occurring 20 to 40 min after oxidation had started.

Jan 1, 1964

INTRODUCTION Trade financing is a particularly important component of short-term financing for a minerals company since errors, extra costs, or payment delays/defaults can easily wipe out the profit margin on a particular shipment or can cause problems when arranging purchases of e.g. reagents. It is required to make any minerals operation function and does not have to be cross-border to require proper management. Sources of working capital are as follows: 1. Extended credit terms from suppliers 2. Inventory loans from banks, finance companies 3. Accounts receivable financing or factoring 4. Credit lines or overdraft facilities from banks. TRADE CREDITS These are provided under the familiar terms "net 10 days" or payment due by the "10th day of the following month." Thirty days for payment is usual but longer periods can be negotiated with regular suppliers. Shipment may be made on open account with specified periods for settlement. In the mining industry, a custom smelter gains trade credit by paying for concentrates three months after the month of shipment. INVENTORY LOANS Various forms of inventory finance are provided via "floor-plan" or warehouse financing. A set percentage, usually 60%-80%, of a good-quality, readily saleable inventory item (e.g. bullion, concentrates, ingots) will be provided by a bank or finance company which will usually have security over these products and has a right to inspect the goods from time to time. Legal security will he achieved by warehouse receipts, floating liens, chattel mortgages, trust receipts, or through collateral management service companies. The latter certify control over stock or distribution outlets and often carry insurance against fraud. Costs range around 1.5% - 2.0% p.a. or more plus some fees to cover out-of-pockets or other costs. ACCOUNTS RECEIVABLES Accounts receivable financing is based on the company's generation of satisfactory receivables which are continuously assigned to a bank or finance company ("factoring"). Because of the risks in being able to collect receivables, such as payments to a smelter for its blister copper shipments to a refinery, this form of financing costs as much as 3% - 5% of the debt amount purchased plus an additional charge for processing paperwork. Receivables can be purchased with or without recourse and on a notification or non-notification basis to the client's customer to, respectively, pay the invoice directly to the bank or not. LETTERS OF CREDIT Letters of Credit ("L/C's") are a common feature of international trade finance whereby a bank, at the request of the buyer, issues a written undertaking to an exporter to pay a specified sum of money within a set time on presentation of previously specified documents. This undertaking means that the exporter is no longer relying on the credit standing and integrity of the foreign importer. A dispute about shipment does not otherwise affect the obligation to pay upon presentation of a "clean" set of documents. These documents include the hill of lading and other documentation, often on standard "form" bases. There are a number of conventions and techniques in L/C's such as irrevocable/revocable L/C's, time L/C's, confirmation/advising L/C's and transferable L/C's. These are tools to adjust for the varying risks that the importer will not pay. Other routes where one is comfortable with an importer include documentary sight hills (D/P hill), documentary term bills (D/A bill), and, of course, open account.

Jan 1, 1985

By Frank C. Smith

THE following article covers the history of a metallurgical campaign, commenced in March, 1905, at the mines of the Socorro Gold Co., in the so-called desert region of Yuma county, Arizona. The results obtained were simply a function of necessity, as will be seen from the description of the local economic conditions and the character of the ores treated. The mines and mill of the Socorro Gold Co. are located near the small village of Harrisburg, about 60 miles from Congress Junction, Arizona, which, at the time this work was undertaken, was the nearest railroad station, and from which all supplies were brought by wagon through the desert, at a cost varying from $10 to $15 per ton in addition to the railroad freights. Wood, the only available fuel, cost $7 per cord. Water for all purposes was pumped to the plant from a well in the valley about 4 miles away. Under the Arizona law, a working-shift is 8 hr. Wages were $6 per day for foremen and mill-men, $4 per day for blacksmiths and engineers, $3.50 per day for miners, and $3 per day for shovelers and trammers. The ore consisted of quartz, in which, above the 250-ft. level, the iron-minerals were largely oxidized and some free gold was visible; below that level few traces of oxidation occurred, and pyrite constituted the principal mineralizer in the quartz, together with occasional pockets of galena and a few eccentric specks of covellite. The 20-stamp mill was equipped with plates for amalgamation, and three Standard tables for concentration. The tails from the tables were elevated by a centrifugal pump to a launder, by which they were conveyed to a 60-ton cyanide-mill for further treatment. The cyanide-mill was erected at so great a distance from the stamp-mill that a separate crew of workers and a separate power-plant were necessary. No arrangement had been made for the treatment of the concen-

Jan 1, 1907

By C. K. Gupta, P. K. Jena

Niobium (columbium) metal in the form of a button has been produced by calciothermic reduction of niobium pentoxide using sulfur as the heat booster. In these experiments with 50 g of niobium pentoxide per batch, the influence of percentage of calcium in excess of the stoichiometric, the percentage of sulfur by weight of the niobium oxide, and the outer-wall temperature of the bomb on the quality and yield of the metal has been studied. The maximum yield on this scale was 78 pct using 50 pct excess Ca and 20 pct of S and at a bomb-wall temperature of 600°C. In experiments conducted on 500 g of niobium pentoxide, reduction with 40 pct excess Ca and 20 pct of S, and at a bomb-wall temperature of 500°C, the.yield of the metal improved to 82 to 84 pct. Some of the reduced-metal samples have been electron-beam melted and the button melts thus obtained have been observed to be extremely ductile and capable of cold reduction to 98 pct without intermediate annealing. NIOBIUM, because of its high melting point (2468oC), its ductility, its resistance to corrosion especially in liquid metals, its high-temperature mechanical properties, and its relatively low capture cross section (1.1 barns) for thermal neutrons, has potentialities as a structural material in the field of nuclear engineering. Niobium and its alloys are also gaining importance in high-temperature technology as components in aircrafts, jet engines, and missiles, electronics, chemical engineering, and surgery. In 1904 Weiss and Aichel 1 produced niobium by reduction of its pentoxide by misch-metal. Since then a large number of investigations have been carried out for the extraction of niobium from its compounds. The various routes through which the metal niobium has been extracted from its compounds can broadely be classified into the following groups: a) reduction of its halides, oxides, and fluo-salts with metals like calcium, magnesium, sodium, and aluminum;2-10 b) reduction of its halides with hydrogen and oxides with carbon or carbides;11-18 c) electrolytic winning from its halides and fluo salts;19-25 and d) thermal decomposition and disproportion ion of its halides26-30 Among the processes in commercial use at present are sodium reduction or fused-salt electrolysis of potassium niobium fluoride, carbide or carbon reduction of the oxide, and hydrogen or metallothermic reduction of the chlorides. In 1907 Von Bolton2 prepared niobium by reducing niobium pentoxide with aluminum. Dennis and Adam-sona attempted reducing the oxide with magnesium powder in argon atmosphere and the resulting metal contained as much as 5 pct O. Mondolfo8 claimed a method for the reduction of the oxides with aluminum. In the case of both magnesium and aluminum reduction, it has been stated31 that by-product oxides such as MgO or A12O3 would involve a major separation problem. Further, in the case of alumino-thermic reduction, the solubility of aluminum in niobium and intermetallic formation have to be reckoned with. In 1922 Bridge3 produced niobium by calcium reduction of niobium pentoxide. Dennis and pdamson' obtained the metal by a similar method with an oxygen content of 0.2 pct. In most of the cases of metallothermic reduction of niobium oxides mentioned above, the metal niobium is obtained in the form of powder. In an attempt to produce massive niobium metal, lock' used iodine as the thermal booster in the bomb reduction of niobium pentoxide by calcium. Massive niobium metal was obtained with a maximum yield of 75 pct containing 0.1 pct 0. Joly32 carried out experiments on calcium reduction of niobium pentoxide on a 2-kg scale using sulfur as the heat booster in a sealed bomb filled with argon. In his experiments, the reaction was initiated by passing a current through niobium spiral wire embedded in the charge. In a typical run consisting of 2000 g of niobium pentoxide, 3100 g of calcium, and 800 g of sulfur (S/Nb2O5 = 3.3), he obtained massive metal in the form of a very uneven mass with a yield of -57 pct. With lower amounts of sulfur and calcium, powder niobium mixed with small metal globules was obtained and the yield was also poor. From these experiments he concluded that this method is not a suitable one for industrial development considering the quality and yield of metal obtainable and the possible hazards with higher ratios of sulfur to niobium pentoxide in the -~charere. In the present work, a detailed investigation on the calciothermic reduction of niobium pentoxide in a stainless-steel bomb in presence of sulfur as the heat booster was undertaken on a laboratory scale.

Jan 1, 1964