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By A. L. Baxley

An Untapped Energy Resource As much as 20 billion cubic feet of natural gas per day are flared from remote oil fields for lack of a commercially viable means of capturing, transporting, and marketing such gas. The magnitude of these gas flares can be put into perspective from an early satellite photograph (Fig. 1) which shows lights from the major cities of Russia and Eastern Europe dwarfed by the natural gas being flared in the Persian Gulf. Together, these wasted resources contain the energy equivalent of about one-half of the gasoline consumption in the United States today (Fig. 2). Additional trillions of cubic feet of natural gas are "shut-in" because of no economically viable means of commercial recovery. Methanol and liquified natural gas (LNG) are the only two practical fuel products which can be produced economically from these gas supplies. Many of these gas supplies are less than 500 million cubic feet of gas per day, making an LNG facility uneconomic. In contrast, barge-mounted methanol plants can economically convert billions of cubic feet of gas per day into safe, clean-burning methanol. The methanol approach offers the only economical route to transform vast, known reserves of natural gas into a highly versatile primary liquid fuel. Methanol Barges: An Innovative Solution The barges will be towed to suitable offshore and upriver locations such as Alaska, South America, Africa, Southeast Asia, Australia, New Zealand, and the South Pacific Islands, as well as fields in the Persian Gulf and Mediterranean Sea. At the offshore production site, a barge will be anchored by a single point mooring buoy that will also serve as an entry point for natural gas feedstock and an offloading point for methanol (Fig. 3). At some sites the barge would be beached. Each barge will produce methanol and store it in internal tanks with a capacity of 18 million gallons. The methanol will be offloaded into conventional tankers and safely transported directly to market. Unlike LNG, methanol requires neither specially built carriers nor specially built receiving terminals. Once a particular gas field has been exhausted, the barge will be towed to another location to continue production. Each barge will measure 320 by 500 ft, approximately the size of four football fields, and will have the capacity to produce 1 million gallons or 2800 metric tons of methanol per day, from approximately 100 million cubic feet of natural gas per day (Fig. 4). The barges will use the highly successful "low- pressure" design developed by the Lurgi Company of Germany, a process proven in land-based methanol plants throughout the world during the last ten years. The decision to use Lurgi technology for "sea-trans- portable" methanol plants was based on the higher efficiency and greater operability of the technology compared to other commercially proven processes. The conversion plant will be designed to accept a wide variety of feed gas compositions, and will produce chemical-grade methanol for the broadest market base (Fig. 5). To minimize costs and construction time, the barge-mounted plants will be built in the high technology environment of a domestic or foreign shipyard. Selection of the construction site for each barge will be dictated by the location of the production site and by the relative construction costs. A number of shipyards have the capacity to build several barges per year. The detailed marine engineering to integrate the design of the processing plant with the floating platform can be performed by numerous major engineering companies around the world. Production Economics The barge-mounted plant concept not only assures large volumes of methanol, but it also keeps the overall production cost low by minimizing construction cost and providing access to low cost natural gas feedstock with no alternative or a negative value. Together, these advantages make the barge-mounted methanol plants economical today. The cost structure of a new barge-mounted methanol plant differs from that of existing methanol producers around the world (Fig. 6). For example, if a U.S. Gulf Coast producer is paying $4.70/MMBtu in 1985 for natural gas, the barge plant could afford to pay about $1.6O/MMBtu for gas and be able to deliver methanol to the Gulf Coast at the same price. At some future date such as 1990, a gas cost of $6.70/MMBtu for a domestic producer would have cost parity with about $3.60/MMBtu gas cost for the barge plant. In many foreign markets, feedstocks other than natural gas are used for methanol production (Fig. 7). For example, most of Japan's capacity is based on LNG while Western Europe uses residual oil or naptha. Because these feedstocks are substantially more ex- pensive than natural gas used by U.S. producers, the barge plants will compete even more favorably in these foreign markets. As crude oil prices rise, the value of methanol in each of these markets will increase. However, the hierarchy of methanol values in these markets should remain unchanged. Furthermore, the cost advantage for using methanol in these markets will improve as world energy costs increase since the value of remote gas should not escalate significantly.

Jan 1, 1982

By Edward Nichols

These deposits, which are found in James County, Tennessee, differ both in their mode of occurrence and in their chemical constitution from any other ores belonging to the same geological horizon which are known to the writer. Taylor's Ridge, or White Oak Mountain, is a long monoclinal ridge, which starts from a point near the Tennessee River in Meigs County, and, following a southwwt .course, extends unbrokenly into Cherokee County, Alabama, where it finally disappears. It is made up of strata of the Upper Silurian rocks, which dip to the southeast. To the eastward of this ridge, which is generally several hundred feet high, are a series of low foothills or knobs, composed of Clinton shales carrying seams of fossiliferous hematites. In the neighborhoud of the East Tennessee, Virginia, and Georgia Railroad, where it crosses the ridges, considerable ore has been milred from these seams during the pst ten years. Most of the workings have been superficial, and the ores mined, with one exception,,carry a silicious and nod-calcareous gangue, and contain from fifty to sixty per ceut. of iron.

Jan 1, 1882

By Charles W. Schwartz, Subash Kolluru

INTRODUCTION Creep of rock in situ, like most rock mass behavior, will be largely governed by the behavior of the natural discontinuities -- bedding planes, faults, and joints, in particular, Several past studies have investigated various aspects of the time-dependent deformation of rock discontinuities; examples include Amadei and Curran (1980), Engelder and Scholz (1976), Johnson (1975), Kaiser and Morgenstern (1979), Solberg et al. (1978), and Wawersik (1974). These studies have provided much insight into the general phenomenon, but they have also produced some inconsistent conclusions and left many unanswered questions. This paper will attempt to shed some light on one of these questions: What is the influence of stress level on the creep of unfilled rock joints? The discussion will follow a two-part approach to the problem: (1) laboratory creep experiments on small scale jointed specimens, (2) a simple theoretical mechanism to explain, at least in qualitative terms, some of the results from the experimental program. EXPERIMENTAL PROGRAM Uniaxial creep tests were conducted on intact and jointed samples of a specially formulated gypsum plaster "synthetic rock." Synthetic rock materials have several advantages over natural rock for long-term creep testing: (a) higher creep rates, (b) lower strength and therefore lower loading requirements, (c) simpler specimen preparation, and (d) more controlled and consistent properties. Gypsum based plasters have been used previously in studies of short-term strength and deformability of jointed samples (Patton, 1966; Einstein and Hirschfeld, 1973). Furthermore, Williams and Elizzi (1977) demonstrated that natural gypsum, at least, creeps in a manner similar to that of most other creep-susceptible natural rocks.

Jan 1, 1982

By B. K. Mozumdar

This paper describes the development and application of a dynamic simulation model of ground movement in an underground mine. Ground movement and related stress distributions with or without stowing are computed as the mining face advances. Pressure profiles plotted for different face positions indicate the presence of a front abutment with a maximum value of four times the original ground pressure but no back abutment in the gob. The surface strain is considerably reduced with increase in packing density in the gob up to a certain value. Beyond this value, the reduction in the strain is minimal, indicating an economic level of packing. As a result of model experiments with different physical systems and geological parameters, it has been concluded that when one of the coordinates for the subsidence and the strain curves is made dimensionless in terms of the critical width, the relationship will hold true for all physical situations. This was in agreement with the NCB observations for different panel lengths.

Jan 1, 1977

By M. Hoch

Equations are derived from statistical considerations to represent the activities of each component of an interstitial solution, and of a compound with a wide homogeneity range as a function of composition and temperature, based on the interaction energies between the various components, Eij. The equations are checked on the binary systems Fe-C, Ta-H, and Y-H, the ternary systems Ti-O-H, liquid Fe-C-0, and the compounds TiO, MnO, COO, and YH,. In all cases the present model represents the experimental data well. THE aim of this paper is to derive from statistical considerations the thermodynamic behavior of non-substitutional solid solutions. "Nonsubstitutional solid solution" means either an interstitial solid solution (carbon in iron, oxygen and hydrogen in hexagonal and bcc titanium, and so forth) or deviation from stoichiornetry in compounds (introduction of vacancies—removal of atoms in one or both lattice sites of compounds FeO, MnO, TiO, NbO, YH2). wagnerl and Fowler and Guggenheim considered the theory of compounds with small deviation from stoichiometry; Anderson includes larger deviations from stoichiometry, the defect being present in one component only, as vacancy and interstitial. Rees considers only vacancies in one of the components and introduces various kinds of sites for the component where the defect occurs; the various sites interact with different energies, thus giving isotherms with two or more plateaus. Both Rees and Anderson are mainly occupied in determining the limits of two-phase regions, and the defect interaction energy is determined from the critical point. Darken and Gurry5 gave a brief treatment of carbon solution in iron; Speiser and spretnak calculated the configurational entropy for binary interstitial solid solutions by assuming that each interstitial atom in the solution excludes a certain number of neighboring interstitial sites from occupancy. Use will be made of this concept later. Hoch et al treated transition metal oxides of the NaCl-type structure, and also oxygen solution in titanium.' The main aim was to determine the behavior of the activity of one of the components as a function of the composition in the single-phase region. In the present paper, the various interaction energies are discussed in detail and the similarity between interstitial solid solutions and deviation from stoichiometry in ionic compounds is shown. Both types of systems have in common that at least on one kind of lattice site only one kind of atom occurs; this gives a fixed frame to the system, permitting a much easier statistical treatment than for the substitu-tional solid solutions. (The theory is also applied to ternary systems.) The difference between an interstitial solution and deviation from stoichiometry can be explained as follows. In an interstitial solution in the pure or standard state the host atom or host structure is present alone and all the interstitial sites which are going to be filled when the interstitial atom is introduced are empty. The interstitial solution is now created by introducing atoms into this empty lattice. In the deviation from stoichiometry, both kinds of lattice sites are already occupied in a special fashion (all lattice sites occupied, or the same fraction of lattice sites empty, or one kind completely filled, the other only two-thirds). Deviation from stoichiometry removes atoms from lattice sites, adds them, or removes atoms from one kind of lattice site and introduces them on the other lattice site. Though both cases (interstitial solution and deviation from stoichiometry) lead almost to the same equations, they are treated separately. The treatment given below corresponds to the "regular solution"10 treatment of substitutional solid solutions. Speiser and spretnak's6 treatment is similar to the athermal solution in the substitutional use. a) INTERSTITIAL SOLID SOLUTIONS Theory. It will be assumed that there are N lat-tice sites occupied by the host atom (denoted by 1). Corresponding to those N lattice sites there are N octahedral and EN tetrahedral interstitial sites. If only one kind of interstitial atom is added, this in-

Jan 1, 1964

Victor G. Hills, Cripple Creek, Colo. (communication to the Secretary): In continuation of the subject of the perform ance of "baby machine-drills," presented by Mr. Thane, I sub mit the following records of work done with these drills in the Cripple Creek district of Colorado. Two years ago, the Anchoria-Leland mine began to use baby drills ; and, to make the experiment more complete, began by purchasing two machines, of the same size, but of different make. The little drills were first used for stoping and up raising, in which they achieved at once an unqualified success. Soon after, they began to be used for the smaller branch-drifts and prospecting cross-cuts, while the large two-man drills were still used for the main drifts. Now they have almost entirely superseded the large drills, since it has been found that they will do any ordinary work in the mine at less cost, both for labor and for powder, per ton of rock broken. The large drills, of course, make more rapid progress; but it is only where time is a specially important factor that they are now used. The Portland mine, where, until within a year, only the large machines were used, is now using some small drills for stoping, and also for drifting. Mr. Thane speaks of having two men to operate a baby-drill; but in this district the baby-machine is always worked by one man without help. Again, in Prof. Christy's postscript, the quoted extract from a letter of Mr. L. T. Seymour, the South African engineer, mentions a " 2 1/2-inch Little Giant," as though that were the smallest machine used in that region. These statements leave some doubt as to just what Mr. Thane means by a " baby-drill." To avoid any such uncertainty as to the present communication, I will say that the large or two-man drill in common use at Cripple Creek has a 3 1/8-inch cylinder; while a few mines employ a 3-inch, and a few, in running large tunnels, a 3 1/2-inch drill. What we call a baby-drill here has a

Jan 1, 1900

By E. Buehler, H. J. Levinstein

The temperature ranges in which the three inter-metallic phases in the Nb-Sn system form have been determined and the composition and structure of two of the three phases has been established. The kinetics of the formation of Nb3Sn in cored wire samples has been studied in the temperature range of 800° to 1050°C. From 800°to 950°C the rate of formation increases by four orders of magnitude. The rate-controlling step for the formation process in this temperature range appears to be the diffilsion of tin through NbSn. At higher temperatu~es a change occurs in the mechanism of the formation process such that up to a temperature of 1050°C the rate of formation of Nb3Sn does not increase above the rate observed at 950°C. For temperatures helow 950°C the current-carrying capacity of the wire increases with increased percent reaction reaching a maximum value when the formation process is 90 to 95 pct complete. The maximum current-carrying capacity obtainable in this temperature range is independent of the temperature. Above 950°C tlze current-carrying capacity obtainable in the wire decreases with increasing temperature of formation. A model is proposed which accounts for the ohserved behavior. RECENTLY, Buehler et a1.l reported the results of an investigation of the process variables which influence the superconducting properties of Nb3Sn-cored wire. These results indicated that at least four variables affect the properties of the manufactured wire. These include composition, particle size of the starting powder mix, temperature of heat treatment, and time of heat treatment. In order to understand completely the role of these variables, it is necessary to have an accurate knowledge of the phase equilibria in the Nb-Sn system. At the present time, phase-equilibrium diagrams for the Nb-Sn system have been published by a number of investigators.2-5 The diagrams differ as to the number of phases present, the composition of the phases, and the temperature range of stability of the phases. The present investigation was undertaken in order to resolve these differences. Since the investigation of Buehler et al. demon- strated that the length of time at the temperature of heat treatment affected the superconducting properties of Nb3Sn, it is apparent that it is necessary to understand the kinetics of the formation process as well as the equilibrium conditions before a complete understanding of the system is possible. As a result, the kinetics of formation of the various phases in the system were also studied in this investigation. EXPEFUMENTAL PROCEDURE Diffusion couples and sintered powdered compacts were employed in the phase-diagram investigation. The diffusion couples were made by filling 1/8-in.-ID monel-sheathed niobium tubes with tin. The monel sheath was employed to facilitate drawing.' The tubes were then drawn to a tin-core diameter of 32 mils. Samples approximately 3 in. long were then cut from the drawn composite. The tin was drilled out of the ends to a depth of 1/4 in. and niobium-wire plugs were inserted into the ends and peened over. The monel was removed by etching in concentrated nitric acid, after which the samples were sealed in evacuated quartz bulbs and heat-treated in a resistance-wound tube furnace. The samples were quenched into ice water upon removal from the furnace. The diffusion couple samples were examined metallographically employing a chemical etching solution consisting of 10 ml of saturated chromic acid per g of NaF. In addition, two anodizing solutions were used for phase-identification purposes. The first was the picklesimer7 solution; the second consisted of equal parts by volume of 30 pct H2O2 and concentrated NH4OH to which 1 g of NaF was added per 25 ml of solution. The anodizing conditions for the second solution were 2 v and 100 ma with a tin cathode. The powdered compacts were made by pressing previously mixed powders of 99.9 pct pure Sn and 99.6 pct pure Nb supplied by the United Mineral Co. into cylinders 3/8 in. in diameter by 1/2 in. long. The cylinders were then sealed in quartz tubes and heat-treated in the same manner as the diffusion couples. The samples were examined metallographically and by X-ray diffraction techniques. Since it was desirable to be able to correlate the kinetic data with current-carrying capacity, the type of specimen chosen for this part of the investigation had to be a compromise between the optimum system for studying kinetics and one which was suitable for making current-carrying capacity

Jan 1, 1964

By I. W. Wark

AN attempt has been made to compare the influences of the two most widely used depressants-alkalis and sodium cyanide-and the most widely used activator-copper sulfate-on the air-mineral contact induced by five sulfur-bearing compounds: (1) sodium di-ethyl dithio-phosphate (sodium Aerofloat); (2) potassium ethyl xanthate; (3) sodium di-ethyl dithiocarbamate; (4) potassium iso-amyl xanthate; (5) potassium di-n-amyl dithiocarbamate. Three different organic types are included in this group and for two of these types a higher and a lower homologue have been tested. The study of potassium ethyl xanthate, the preliminary results of which have been described,1 was continued until there had been obtained a comprehensive account of the influence of variations in pH value and in the concentrations of xanthate, sodium cyanide and copper sulfate on the adsorption of this xanthate by galena, sphalerite, pyrite and chalcopyrite. With concentrations of copper sulfate and collector that the work with ethyl xanthate indicated would give the most illuminating results, the other four collectors were tested in a less exhaustive way. The same general principles determine the influence of alkalis, cyanide and copper sulfate, no matter to which type the collector belongs, but there is a gradation in the quantitative measure of such influence. In the list above, the five collectors are cited in the order of their capacity to withstand the depressants, sodium Aerofloat being the least resistant, potassium di-n-amyl dithiocarbamate the most resistant. This order coincides substantially with that of the solubilities of the heavy metal salts formed by double decomposition with the collectors, those formed from sodium Aerofloat being most soluble. There is little difference between the third and fourth members of the series, and under certain conditions their positions are reversed.

Jan 1, 1934

By G. P. Bowman, D. F. Howard, R. K. Saxer, J. R. Myers

This study was undertaken to redetermine the phase diagram of the lead molybdate-bismuth molyb-date salt system. A portion of this diagram was originally determined by Zambonini in 1920, and more recently by Beale. Thermal analysis and X-ray methods were used to define this system. The diagram was found to be a simple eutectic type with an eutectic at 11 mol pet lead molybdate and 634OC. A maximum solid solubility of 6 mol pet was found at the lead molybdate end of the diagram. Solid solubility at the bismuth molybdate end of' the diagram appeared to be quite limited and possibly nonexistent. These results generally agree with those of Beale, but differ significantly from those of Zambonini. MOLTEN salt systems have been vigorously investigated since World War 11. Much of this work has been performed as pure research; however, consideration has recently been given to their possible use as coolants in nuclear reactors. Because of this application and other potential applications, it was decided to reinvestigate the lead molybdate-bismuth molybdate phase diagram. This diagram was originally determined by Professor F. Zambonini of the University of Turin in 1920.' In 1961, a portion of this diagram was redetermined by Beale.2 With the quality equipment and high-purity materials available today, it was expected that discrepancies between the results of this present investigation and Zambonini's work would be obtained. This investigation was conducted in two parts. First, the general shape of the phase diagram was determined by cooling-curve analysis. Next, the primary solid-solution phase at the lead molybdate end of the diagram was investigated by means of X-ray diffraction techniques. EXPERIMENTAL PROCEDURE High-purity lead molybdate, assaying 99.993 pet, and high-purity bismuth molybdate, assaying 99.989 pet, were employed during this investigation.* The results of a spectrographic analysis of the two salts are presented in Table I. Twenty-two experimental compositions were prepared for cooling-curve analysis. Compositions selected are presented in Table 11. Several randomly selected samples were analyzed chemically to check that the compositions of the prepared specimens were within the limit of

Jan 1, 1964

By F. W. Warnaars

The Los Bronces copper deposit is located on the west side of the Andes Mountains in central Chile about 54 km (34 miles) northeast of Santiago. The deposit consists of a hydrothermal breccia complex that was emplaced in the western part of an earlier porphyry copper mineral system. This system exhibits propylitic, sericitic, silicic to potassic hydrothermal alteration with disseminated and stockwork copper-molybdenum sulfide mineralization over an area of about 12 km2 (5 sq. miles). The Rio Blanco Mine, about 800 m (0.5 miles) east of Los Bronces, is an underground block caving mine located in the central part of this system. The mineralization at Rio Blanco is characterized by one of the highest primary copper grades known for porphyry copper systems. Present mining at Los Bronces is from an open pit located in the northern part of a hydrothermal complex that is composed of at least seven different copper-bearing tourmaline breccia pipes. Together these pipe form one large contiguous body about 2 km (1.2 miles) long and 0.7 km (0.4 miles) wide at the surface. This body crops out at elevations of between 3.5 and 4.1 km (2 and 2.5 miles) above sea level. Each breccia pipe has a characteristic shape, matrix, type and degree of mineralization and alteration. Quartz, tourmaline, specularite, anhydrite, pyrite, chalcopyrite, sericite, chlorite, and rock flour are present in significant quantities in most breccia matrices. The clasts consist mainly of quartz monzonite, andesite, and minor amounts of quartz latite porphyry. The seven different breccia pipes are identified, from oldest to youngest, as Ghost, Central, Western, Infiernillo, Fine Gray, Anhydrite, and Donoso. The breccias coalesce, interfinger, or display gradational contacts. In places they form sharp contacts with each other and with the surrounding intrusive or volcanic host rocks. Early breccias are disrupted or cut by subsequent breccia events. Explosive emplacement was followed by collapse after the release of hydrothermal pressures. The Los Bronces breccia complex has sharp contacts with the surrounding intrusive stock and andesites. These contacts dip inward on the north, west, and south margins. The eastern contact is nearly vertical and/or dips slightly outward and suggests a westward tilting of the body of about 15° after emplacement. The breccia complex is not fully explored at depth, but is known to extend at least below the 3 km (2 mile) elevation in its central eastern part, about 1 km (0.6 miles) below the highest breccia outcrop. Primary copper sulfide minerals are most abundant in the Donoso and Infiernillo breccias where the matrix consists mostly of tourmaline, quartz, sulfides, specularite, and/or chlorite with minor amounts of rock flour, sericite, and clay. The volume of the matrix material is mostly smaller than in the other breccias. Secondary enrichment enhanced the primary copper grade in the southern two-thirds of the Los Bronces breccia complex and in much of the surrounding porphyry system. Primary grades of 0.3% to 0.5% copper have been enriched to approximately 0.9%. This enrichment may extend from surface outcrops to a depth of more than 500 m (1, 640 ft) in more permeable parts of the breccia complex. A leached and oxidized capping of variable thickness covers a large part of the enriched layer and exceeds 300 m (984 ft) at one location. The shape and depth of the enrichment blanket and the capping suggest that the enrichment process is still active.

Jan 1, 1983

By G. St. J. Perrott

ONE of the important hazards in coal mining is the danger of ignition of explosive mixtures of methane and air or coal dust and air, or both, by the explosives used in blasting the coal. It has long been recognized that certain explosives are more dangerous than others in this respect. In America and abroad, government testing stations have been established for officially passing on the suitability of explosives for use in coal mines. In this country the Federal government is given no power to enforce the use of the approved explosives but merely publishes from time to time, through the Bureau of Mines, those explosives which have passed the tests necessary to place them on the list for use in coal mines. In spite of this lack of police power, the increase in the use of permissible explosives since the promulgation of the first list in 1909 has been rapid. In 1912, only 8 per cent. of the explosives used in coal mines were permissible; in 1925, this figure had increased to 25 per cent. While the past record of permissible explosives has been most gratifying, it is very desirable that we pave the way for further improvement in explosives and blasting methods by learning more about the mechanism of ignition of gas and coal dust by explosives. To this end a considerable amount of investigative work already has been carried out and a large amount yet remains to be done. It is the purpose of this paper to discuss briefly the effect of the characteristics of explosives and method of loading in the borehole on the liability of a blown-out shot to ignite explosive mixtures of gas or dust. For an account of the methods by which an explosive is tested for permissibility for use in coal mines, the reader is referred to Schedule 17-A of the U. S. Bureau of Mines.

Jan 10, 1926

By Louis V. Bender

AfteR investigating the work of coal-dust fired reverberatories of the Canadian Copper Co., at Copper Cliff, Ontario, the management of the Washoe Reduction Works decided to experiment with and ascertain the advantages of using coal dust as a fuel in their reverberatories. Consequently, during the month of June, 1914, reverberatory furnace No. 8 was changed in order to use powdered coal as a fuel. The results obtained by this method of firing are gratifying and show a decided saving in cost of smelting as compared to grate firing with lump coal. The furnace, as remodeled, is 124 ft. long by 21 ft. wide, and varies in height from 8 ft. 6 in. at the back end to 5 ft. 7 in. at the skimming end. The general construction of the furnace is similar to that of the other furnaces at this plant. There are no side doors to this furnace, as it was thought that with the present arrangement for feeding no "fettling" or "claying" would be required. The interior of the furnace can be inspected through the burner port holes, after shutting off the burners and giving a few seconds' time for the gases inside the furnace to clear away. The charging is done on either side of the furnace from longitudinal hoppers, extending a distance of 74 ft. from the back end of the furnace. Leading from the hoppers into the furnace are 6-in. pipes, spaced 191/4 in. apart, through which the charge is intermittently dropped. The charge is kept well above the slag line at all times; in this way the side walls are protected and no fettling is needed on this portion of the furnace. The remainder of the furnace requires fettling. After operating for three months, we found that the bricks were eaten into along each side wall from the skimming door back to the point where the charge had been dropped. The depth of this cutting away was 8 in. close to the front end and gradually tapered to zero at a distance of 50 ft., and was greatest on the side of the furnace having the larger flue connection. Hoppers will be put in for the entire length of furnace, from which fettling material will be dropped, to prevent this cutting. After the run of three months, the roof was in excellent condition. At the back of the furnace the bricks were not cut into at all; at 30 ft.

Jan 1, 1915

By R. W. Raymond

ChaRles William HenRy Kirchhoff was born March 28, 1853, at San Francisco, Cal., where his father, Charles Kirchhoff, was at that time consul for his native country, Germany. A few years later, the family moved' to Hoboken, K. J., in which city the son received his preliminary education at the Hoboken Academy, proceeding later to Germany, where he was graduated in 1874 as mining engineer and metallurgist at the Prussian Royal Mining Academy of Clausthal, in the Harz. Upon his return to the United States, he became chemist of the Delaware Lead Refinery at Philadelphia, and retained that position for three years. But in 1877 he began what was to be an almost uninterrupted life-long association with David Williams, the publisher of The Iron Age, of New York, who established in that year, and continued for a brief period, a journal entitled the Metallurgical Review, on the editorial staff of which Mr. Kirchhoff received a place. The enterprise was doubtless an attempt to cover a wide metallurgical field outside of that which properly belonged to The Iron Age. But it was soon abandoned, and Mr. Williams wisely concentrated his energies upon the older journal, which, under his vigorous and skillful management, and the labors of the able editors and correspondents whom he selected, became one of the greatest institutions of its class in the world. From 1878 to 1881, Mr. Kirchhoff was assistant editor of The Iron Age. From 1881 to 1884, he was managing editor of the Engineering and Mining Journal; in 1884 he returned to The Iron Age, to become for five years its associate editor, and in 1889, on the retirement of James C. Bayles, editor-in-chief. This list of dates and employments, without further comment, sufficiently indicates, at least to the eye of an expert, that Mr. Kirchhoff had found his congenial career in trade journalism. His qualifications for this profession were somewhat exceptional. He possessed the scientific training, the knowledge of foreign languages and literatures, and the power of making and keeping friends, which enabled him to get early notice of technical novelties; he had the taste for statistics which made him both industrious and intelligent in their collection and use; and to these traits he added a mastery of the meaning of such accumulated data, and a sane, critical judgment of the situations which they represented, as well as of the sources and the figures themselves, which made his opinion weighty

Jan 1, 1917

THE Pacific Northwest's joint meeting of the Industrial Minerals and Metals Branch in Spokane, Wash., drew 260 members, 72 students, and 30 ladies. Theme of the meeting was to acquaint men in the industry with students and younger men. The special appeal to the students was coupled with an attempt to attract young technical men from the companies located in the northwestern area. The activities of the meeting were available without charge to students. Contributions were made by the industries in the area to assist student attendance. Following the special student and faculty trip to the Trentwood rolling mills and research laboratory on Friday afternoon everyone enjoyed a complimentary dinner. A. Y. Bethune of the Sullivan Mining Co., spoke to the group on the operations of Kellogg, Idaho electrolytic zinc plant. The Industrial Minerals Div. session was opened on Saturday morning by R. G. Vervaeke, Conference Chairman who introduced A. O. Bartell of the Raw Materials Survey, Portland, Ore. Mr. Bartell in turn introduced C. W. Sweetwood who discussed Open-Pit Phosphate Mining.

Jan 1, 1952

By Gardner F. Williams

These mines are situated in Griqualand West, which is now a part of the Cape Colony. The town of Kimberley is in latitude 28o40' South, longitude 25o10' East, about 640 miles northeast of Cape Town and 450 miles from Port Elizabeth on the east coast. It is connected with these towns by a railroad, which has been completed during the present year. The mines are located in a desert country, resembling the desert portion of Arizona. The elevation above sea-level is about 4000 feet. The climate is exceedingly hot during the summer months. Diamond Deposits.—There are four large diamond-bearing deposits at Kimberley, all lying within a radius of one and one-half miles, having the De Beers mine as the center. These are named the " Kimberley," " De Beers," " Du Toit's Pan," and " Bultfontein " mines. There is, also, directly west of the Kimberley mine, a small deposit, called " St. Augustiue," which has been prospected only, and

Jan 1, 1887

By D. E. Spiller

Introduction Precious metals recovery from ore can generally be accomplished using gravity concentration, flotation, and/or hydrometallurgical (leaching) techniques. The objective of this paper is to show why gravity concentration can be an important part of recovery systems that employ leaching as the primary unit operation. A brief discussion of modem gravity concentration equipment is also presented. Discussion Gravity concentration of ores has generated increasing interest in recent years. Reasons for this interest include: • Gravity concentration is environmentally attractive. There is little or no use of reagents. Hence, it is relatively nonpolluting. • The cost of cyanide has continued to increase. Therefore, cost savings may be realized whenever leaching feed tonnage can be reduced by preconcentration. • Compared to flotation and leaching, gravity equipment costs are low per processed ton. Field installation costs for gravity circuits usually are less because many' units are supplied as self-contained modules. Also, the cost required to supply services, particularly power, to a gravity plant site are also less. In situations where preconcentration at coarse particle size is applicable, significant grinding equipment savings may be possible. • Gravity circuit operating costs are also relatively low compared to typical flotation and leaching circuits. Reagents, power, maintenance, and manpower savings in a well-engineered gravity plant may be realized. Again, if grinding is reduced, significant power and steel (media and liners) savings are possible. •In recent years, more efficient gravity concentrating devices have been developed. Benefits to Precious Metal Leaching Gravity beneficiation can complement precious metal leaching in two ways. First, the recovery of coarse liberated values before leaching may reduce leach time requirements and may reduce reagent consumption. Second, gravity preconcentration can reduce the size of a leach plant by decreasing the quantity of material to be leached. Coarse gold and silver have been shown to leach rather slowly. Kameda (1949) and Habashi (1967) have investigated the kinetics of cyanide leaching systems. They concur that in a heterogeneous reaction, the rate of gold and silver dissolution is directly proportional to the surface area. Thus, the instantaneous rate of dissolution for spherical 0.37 mm (400 mesh) gold is theoretically -25 times faster than for the same amount of gold at .841 mm (20 mesh), based on data from Fuerstenau, Chander, and Abouzeid (1979). Conversely, coarse liberated, +.841 mm (20 mesh), gold is more readily recovered by gravity concentration than is fine, -.037 mm (400 mesh) gold. Therefore, it is apparent that the two recovery systems complement one another. Figure 1 data demonstrates the potential synergism. A sample of - 3.327 mm (6 mesh) Nevada gold-bearing ore was cyanide leached using conventional bottle-roll test procedures. Gold extraction was determined as a function of leaching time. A second sample split from the same leach feed material was hand jigged to remove a coarse heavy mineral fraction, including virtually all of the +.210 mm (65 mesh) liberated free gold. This second sample, with the coarse gold and heavy minerals removed, was subjected to an identical cyanide leach procedure. Figure 1 presents the resulting comparative extraction data. Note that the percent gold extraction for the sample containing no +.210 mm (65 mesh) free gold includes the coarse gold recovered by gravity. The data show that the sample containing coarse gold required about 72 hours of leaching time to achieve 80% extraction. This compared to about 22 hours of leaching time for 80% gold recovery from the sample that contained only -.210 mm (65 mesh) free gold. Thus, there was a 69% reduction in leaching time. The improved extraction data is not wholly attributable to coarse gold removal, but rather it was the combination of gold removal and rejection of other heavy mineral cyanide consumers or leach retardants. Further investigation was not warranted at this time. Preconcentration is the second manner in which leaching systems can benefit from gravity concentration. The premise is that preconcentration can reduce the quantity of leach feed, which, in turn, may reduce leaching costs. Figure 2 presents preliminary data developed by CSMRI for US Minerals Exploration (USMX). Centennial Exploration Inc., in agreement with USMX, is proceeding with evaluations to determine the suitability of various processing schemes for recovery of gold values from the Montana Tunnels property. The data shows how the ore can be preconcentrated by gravity techniques to result in a reduced feed tonnage to secondary extraction techniques, presumably flotation or cyanide leaching. Testing has shown that Reichert cones, followed by treating the cone concentrate on spirals, can deliver about 88% gold recovery in about 13% weight, that is, 87% weight rejection. Consequently, fine grinding and reagent costs are attributable to only 13% of the plant feed rate. Cost data is not yet available, but the potential exists for significant cost savings.

Jan 1, 1985

By P. J. Potter

PRACTICALLY all tin-base babbitt metals used in engine bearings are made to customers' specifications, which are many and varied. The copper ranges from 3 to 8 per cent. and the antimony from 4 to 13 per cent.; generally, the babbitt with lower copper content will contain from 4 to 8 per cent. antimony and that with higher copper will have from 7 to 13 per cent. antimony. The allowable lead content varies from 0.20 to 2.00 per cent. If a solder is used as a bonding material instead of tin, the resulting material in the finished bearing will have a higher lead content, but it would not be enough to bring the lead above specificational limits. Impurities such as iron, arsenic and bismuth should be determined in the tin before using to insure a uniform and high-grade product. These points must be considered when segregating and grading the waste that comes from the finishing operations on lined bearings.

Jan 1, 1929

By H. H. West

Modern mining is vitally concerned with the prompt, efficient movement of great quantities of bulk materials. It is not surprising, therefore, that intensive efforts are being directed toward improving effectiveness and reducing costs of materials handling systems. Historically, the operation of removing bulk material from a moving conveyor and depositing it in utilization or storage areas has been a high cost segment of the materials handling system. Since this operation is often performed by belt trippers, considerable attention has been focused on improvement of these devices. Frequently, the greatest opportunity for improving performance of these devices is the incorporation of remote or automatic control systems. This article describes tripper control systems developed by Kennecott Copper Corp. and explains the control philosophies used in designing these systems.

Jan 3, 1963

By T. Y. Yan

SUMMARY Laboratory high pressure column tests have shown that the presence of 1-20 ppm of aluminum ion effectively prevents permeability loss during uranium leaching with leachates containing sodium carbonate. If added after permeability loss has occurred, aluminum ion can restore the permeability to nearly its original value. No deleterious effect was observed on uranium leaching performance and the technique should be quite compatible with all field operations. INTRODUCTION The recovery of uranium values from underground deposits by in situ leaching or solution mining has become economically viable and competitive with conventional open pit or underground mining/milling systems (Merrit, 1971). In situ leaching processes are particularly suitable for small, low-grade deposits located in deep formations and dispersed in many thin layers. Many such ore bodies occur along a broad band of the Gulf Coastal Plain (Eargle et. al., 1971). The advantages of the in situ leaching processes have been reviewed (Anderson and Ritchi, 1968). In the in situ leaching process, a lixiviant containing the leaching chemicals is injected into the subterranean deposit and solubilizes uranium as it traverses the ore body. The pregnant lixiviant or leachate is produced from the production well and is then treated to recover the uranium. The resulting barren solution is made up with the leaching chemical to form lixiviant for re-injection. Upon completion of the leaching operation, the formation is contaminated with leaching chemicals and other species made soluble in the leaching operation and has to be treated to reduce the concentration of these contaminants in the ground water to levels acceptable to the regulatory agencies (Witlington and Taylor, 1978). Restoration is accomplished by injecting a restoration fluid, which could be the fresh water or water containing chemicals, into the formation. As it traverses the leached formation, the restoration fluid picks up the contaminants and is then produced at the production well. This produced water is either disposed or purified for recycle. In both phases of operation, formation permeability or well injectivity is one of the most important parameters which determines the viability of the in situ leaching process. Low formation permeability limits production rates, leading to uneconomical operations. The formation is said to be sensitive if there is a sharp loss of permeability on contact with water and other fluids. Many uranium bearing formations, for example, the Catahoula formation of the Texas Coastal Plain, contain significant amounts of clay minerals which are water sensitive. Serious permeability losses can occur when the pH and chemical composition of the lixiviant is significantly different from that of the formation water. Jones has investigated the influence of chemical composition of water on clay blocking of permeability (Jones, 1964) and Mungan studied permeability reduction through changes in pH and salinity of the water (Mungan, 1965). Various mechanisms of permeability damage have been proposed and reviewed (Jones, 1964; Mungan, 1965; Gray and Rex, 1966; and Veley, 1969). When large amounts of swelling clays are present, a significant fraction of the flow channels in the formation can be reduced due to swelling. However, in most cases, swelling need not be the main cause of permeability losses. Particle dispersion and migration or clay sliming can be more important causes for formation damage. Clay particles entrained in the moving fluids are carried downstream until they lodge in pore constrictions. As a result, microscopic filter cakes are formed by these obstructions, plugging the pores, effectively restricting fluid flow and reducing the formation permeability. Moore found that as little as 1-4 percent clays present in a fine grained sandstone could completely plug the formation if they are contacted by incompatible injected fluids (Moore, 1960). It has been found that injection of NaHC03/Na2CO3 lixiviant into formations with significant clay content often leads to loss of formation permeability and well injectivity. To alleviate this problem a change of the lixiviant composition to KHC03/K2CO3 has been proposed. At present, however, many in situ leaching operations employ NH4HC03/(NH4)2C03 mixtures as a source of carbonates. This approach has been successfully used in South Texas by Mobil, Intercontinental Energy, Wyoming Minerals and U.S. Steel, etc. The use of ammonium carbonates solutions, however, contaminates the formation and requires a time-consuming restoration operation. The other approach to reduce the permeability loss is to pretreat the sensitive formation with chemicals which prevent clay dispersion and migration. Such chemicals include hydroxy-aluminum (Reed, 1972 and Coppel et. al., 1973), hydrolyzable zirconium salts (Peters and Stout, 1977), hydrolyzable metal ions in general (Veley, 1969) and polyelectrolyte polymers (Anonymous). Still another approach, is to minimize the "shock" caused by sudden injection by gradually changing the chemical composition of the injected fluids from that of the formation water. THE APPROACH Since permeability loss can be an important factor limiting the efficiency and economic viability of the in situ leaching process, a study was initiated on

Jan 1, 1982

By J. G. Faller, L. P. Skolnick

The distribution of cobalt and vanadium over non-equivalent crystallographic sites in C14-type Zr(VCo), alloys has been investigated. An anomalous X-ray scattering technique developed by Skolnick, Kondo. and lavine7 by which the separation in the scattering factors of two similar atoms can be enhanced was employed. Six alloys spanning the pseudobinary section ZrV1.6Co0.4-ZrVO.6CO1.4 at 10pct steps showed a nonrandom compositionally dependent distribution. Specifically, at high vanadium content cobalt preferentially occupied sites of type (6h) and vanadium, sites of type (2a; at low vanadium content the reverse was observed. In addition to the distribution fraction the structural parameters x and z were obtained. There was no significant deviation of these parameters from those obtained in the ideal C14 structure. Certain suggestions are made to account for the observed nonrandomness in the distribution of atoms on the two types of sites. INTERMETALLIC compounds of formula AB2 iso-morphous with MgCu2, MgZn2, and MgNi2 are known as Laves phases. Because Laves phases exhibit high symmetry and coordination numbers, the highest possible for an AB2-type compound,1 they are among the most frequently observed compounds in nature. In recent years interest has centered about the purely transition metal Laves phases2-' in efforts to understand the function of atomic size and electronic structure in the formation of these compounds. It has been observed that pseudobinary Laves phase systems often show a variation of structure across the phase diagram. Such a system is the ZrV2-ZrCo2 in which the structure varies from cubic MgCu2 to hexagonal MgZn2 to cubic MgCu2.4 Some understanding about the conditions under which the second modification is stable can perhaps be gained by studying the distribution of cobalt and vanadium atoms on lattice sites in the MgZn2 modification of the system ZrV2-ZrCo2. In both the MgZn2 and MgNi2-types there exist nonequivalent positions open to occupancy by the B element, whereas in the MgCu2 prototype all sites are equivalent. Skolnick, Kondo, and La-vine7 have developed an anomalous scattering technique suitable for this type of investigation. Whereas the influence of size on the formation of a Laves phase is well recognized, no substantial evidence has been put forth in support of the size ratio dependence of a particular prototype. Berry and Raynor8 suggested that RA /RB ratio does indeed affect the type of structure that is chosen, MgZn2 compounds tending to cluster about 1.225 while MgCu2 compounds were found at larger deviations from this ratio. Dwight,3 however, from a study of 164 Laves phases does not believe this generalization to be justified. Electronic effects are certain to play a part in the stability of Laves phases in general and in the choice of a structure type in particular. For example, size along would favor the formation of Laves compounds of Ti, Zr, Hf, Ta, or Nb as the A element with nickel or copper as the B element. The absence of such is attributed to an unfavorable electron : atom ratio by Elliott and rostoker.4 Early experiments of Laves and witte9 with pseudobinary and pseudoternary systems of the three prototypes established the dependence of crystal structure upon electron: atom ratios. They observed that the MgCu2 structure dissolved elements of higher valency until the electron: atom ratio of =1.8 was reached; the MgZn2 likewise dissolved elements of lower valency replacing zinc. witte,6 from calculations of the electron : atom volumes of Brillouin Zones, obtained limits of stability for the prototypes MgCu2 and MgZn2. Elliott and Rostoker4 used these limits with considerable success in the all-transition element Laves phases they investigated. According to witte,6 compounds between the electron :atom ratios of 1.80 and 2.32 were of the MgZn2 type; those above and below exhibited the MgCu2-type structure. On the basis of these limits and an assumed valency of zirconium based upon the near tetra-valence of titanium, Elliott and Rostoker obtained valencies for the first transition series elements. For the Laves phases with which this investigation is concerned, ZrV2 and ZrCo2, the authors calculated electron :atom ratios of 2.54 and 1.56, respectively. These ratios are for the MgCu2-type structure and straddle the stability band of the MgZn2 modification. One could, therefore, predict that a pseudobinary system ZrV2-ZrCo2 should pass through the MgZn2 modification in traversing the composition diagram from one end to the other. Implicit in this assumption is a smooth change of the electron: atom ratio from 2'54 to 1.56. MOSS10 states that his finding the low temperature structure of ZrCr2 to be C15 instead of C14 alters greatly Elliott's valency of zirconium and hence the assumed valencies of the other metals. Such a quantitative correlation of structure with electron : atom

Jan 1, 1963