Search Documents

By James A. Barr

The writer's first experience with flotation was during World War I, in the bene-ficiation of Alabama graphite schist ores. One plant used a cone with a peripheral overflow; dried ore was distributed by a rotating disk onto the surface of the water. No reagents were added' The flake-graphite film floated along with some fine silica. Dewatering screens separated the flake. Naturally unit capacities were low but results were satisfactory. Minerals separation mechanical flotation machines were used in another graphite mill. pine oil was added directly to the head cells, with very good results. All the graphite concentrates from both types of flotation were carefully refined after drying by passing through burr mills, followed by very close sizing on sifters. In March 1927, the writer ran a very successful flotation test on minus 28-mesh, deslimed, Tennessee brown phosphate "sand,"† using a mechanical laboratory machine built in the shops of the Anaconda Copper Mining Co., through the courtesy of Ernest Klepetko. The machine was soon converted to a subaeration type, with improved results. This phosphate research program was greatly enlarged and led to the formation of the phosphate Recovery corporation. The development of nonmetallic flotation followed rapidly. The first flotation machine for bene- ficiation of phosphate was installed in January 1928, in a pilot mill, by International Minerals and Chemical Corporation, at Mulberry, Florida. These machines were Minerals Separation, 24-in., sub-,,,tion types, with weir control of pulp between cells. The sand bleeders were very small—about 1¼-in. diameter. As the feed flake. Is to 25 per cent plus 28-mesh material, the cells choked up frequently. This was corrected by cutting a 2 by 6-in. bleeder opening above the impellers, at grid level, which short-circuited the coarse fractions directly into the feed of the next cell. The matter of size classification and grinding was a considerable problem, which does not come within the scope of this article. Because of the troubles noted above, cascade-type roughers were installed in the second flotation plant, built about 1930. Experimental cells had been tried in 1928. General plant layout is shown in Fig. 1 and details of the cascade construction in Fig. 2. The cascade installation in plant NO. 2 gave good results and eliminated troubles due to choke-ups and coarse feed. The capacities and metallurgical results equaled those of the modified 24-in. Minerals Separation flotation machines in plant No. I The objections to the cascade installation are the floor space and mill height required. In 1939 an improved Minerals Separation machine was tested in parallel with the cascades. Extensive tests indicated

Jan 1, 1947

By David E. Nicholas

INTRODUCTION In this paper, a numerical process for selecting a mining method, with the emphasis on underground mass mining techniques, such as caving, induced caving, and stoping, is proposed. In the past, selection of a mining method for a new property was based primarily on operating experience at similar type deposits and on methods already in use in the district of the deposit. Then, the chosen method was modified during the early years of mining as ground conditions and ore character were bet- ter understood. Today, however, the large capital investment required to open a new mine or change an existing mining system make it imperative that the mining methods examined during the feasibility studies and the method actually selected have a high probability of attaining the projected production rates. Although experience and engineering judgment still provide major input into the selection of a mining method, subtle differences in the characteristics of each deposit, which may affect the method chosen or the mine design, can usually be perceived only through analysis of measured characteristics. The parameters that must be examined when choosing a mining method include: 1) geometry and grade distribution of the deposit; 2) rock mass strength for the ore zone, the hanging wall, and the footwall; 3) mining costs and capitalization requirements; 4) mining rate; 5) type and availability of labor; 6) environmental concerns; and 7) other site-specific considerations. This paper encompasses a detailed look at the first two parameters since they, plus mining costs, have the greatest impact on the selection of a mining method. The proposed method selection process is for a project where drilling has defined sufficient geologic reserves, but little or no underground development has been done. Since each deposit has its own characteristic geanetry/grade distribution, and rock mechanics properties, mining method selection should be at least a two-stage process. In Stage 1, the deposit is described in terms of geometry, grade distribution, and rock mechanics properties. Using these parameters, the mining methods can be ranked to determine which are most applicable; they can then be considered in general terms of mining and capitalization cost, mining rate, type and availability of personnel, environmental concerns, and other site-specific considerations. In Stage 2, the most likely mining methods are costed out, based on a general mine plan. Mining and capitalization costs are used to determine a cut-off grade fran which a minable reserve can be calculated; economic comparisons can then be made to determine the optimum mining method and economic feasibility. During the mine planning phase of Stage 2, rock mechanics information would be used to provide realistic estimates of underground opening size, amount of support, orientation of openings, and caving characteristics, and open pit slope angles. If ground control or operational problems should be encountered with the methods being considered, modifications could be made. Although planning on paper extends start-up time, it is cheaper to err on paper than to find the error after mining has begun. METHOD SELECTION - STAGE 1 The main purpose of Stage 1 is to select those mining methods which should be considered in greater detail. The simplest way to do this is by defining those characteristics required for each mining method and then determining whether the characteristics of the deposit are suitable. However, no one mining method is so restrictive that it can be used for only one set of characteristics, as indicated by the classification system proposed by Boshkov and Wright (1973). In the mining method selection proposed, geometry, grade distribution, and rock mechanics characteristics are ranked according to their acceptability for ten general mining methods.

Jan 1, 1981

By A. L. Barnes, A. M. Rowe

A heat transfer study was made of hot gas injection into oil shale through wells interconnected by vertical fractures. This analysis involved the simultaneous numerical solution of a nonlinear, second-order partial differential equation that describes two-dimensional conduction heat transfer in oil shale and a nonlinear first-order partial differential equation that describes convection heat transfer in the fractures. Three nonlinear, temperature-dependent coefficients were used in this work; they are thermal conductivity, thermal capacity and retorting endothermic heat losses of oil shale. Vertical fractures were considered to be of finite height. Although vertical conduction heat transfer was not considered, an estimate of the error resulting from this limitation was made. The effects of injected gas temperature, injection rate, system geometry, cyclic injection and time upon retorting efficiency were investigated. Results from this study show that the rate of retorting oil shale is a direct function of both injection temperature and rate, and the theoretical producing air-oil ratio:(AOR) is an inverse function of temperature. Retorting rates are constant until "breakthrough" of the 700 F isotherm at the producing well, assuming constant injection parameters. Retorting rates for bounded systems are higher than the analogous unbounded systems and likewise AOR's are less. The use of an alternating injection-soak routine with high injection rates is less efficient than continuous injection at lower rates. These results indicate that injection temperatures on the order of 2000F or greater may give theoretical AOR's in the economic range. INTRODUCTION Over half of the known oil shale reserves are located in the U.S., and most of them lie in the Piceance Creek basin of Western Colorado. The Colorado oil shale outcrops on the edges of the Piceance Creek Basin. At the outcrops the shale beds are relatively thin, from 25 to 50 ft thick. In the center of the basin the oil shale is as great as 2,000 ft thick and is covered with 1,000 ft of overburden. It has been estimated that there are over 1,000 billion bbl of oil in shales having an oil content over 15 gal/ton in this basin. Oil shale does not contain free oil but an organic matter called kerogen. Kerogen yields petroleum hydrocarbons by destructive distillation. It must be heated to approximately 700F, at which temperature it decomposes into shale oil, gases and coke. The U.S. Bureau of Mines and, more recently, oil companies have conducted considerable research on surface retorting methods to economically recover oil from this shale. Another approach to exploit the oil shale deposits, in particular that portion having 1,000 ft of overburden, is to retort the oil shale in place and produce the liquid and gaseous hydrocarbons through wells drilled into the shale. Some research has been done on this approach.l,2 There are several variations to the in situ retorting approach. These variations fall into one of two groups, depending upon the geometry of the system: (1) retorting in a highly fractured or broken up matrix; (2) retorting from single fractures between production and injection wells. The latter is the group studied. Several investigators,3-6 using various assumptions, have studied flow of heat through horizontal systems. The objective of this work was to make a heat transfer study of in situ retorting oil shale by hot gas injection through wells interconnected by single vertical fractures of finite height. The oil shale thermal conductivity, thermal capacity and retorting endothermic heat losses were considered to be functions of temperature. A knowledge of the

Jan 1, 1969

By H. C. Chang, N. J. Grant, A. R. Chaudhuri

During the creep of coarse-grained polycrystalline magnesium at elevated temperatures, a nonbasal type of slip was found to play an important role in the deformation processes. The nonbasal slip traces were examined metallographically in eight specimens (13 grains) and the observed glide plane located stereographically for each grain. The tests were run at 500&apos; and 700°F at stresses of 148 to 786 psi. Based on these measurements and theoretical calculations, the crystallographic elements for nonbasal slip were determined. WHEN a metal deforms by slip, the conventional appearance of the slip bands on the surface of a specimen is that of straight lines. Such is the behavior of the face-centered-cubic metals, as for example, aluminum. The particular slip system on which slip occurs in these metals is governed by the criterion of resolved shear stress. Work on the body-centered-cubic metals a-iron,&apos;. &apos; molybdenum, and columbium4 has shown that the process of slip is somewhat more complicated than the simple gliding of one close-packed plane of atoms over another in a close-packed direction and that it results in the formation of wavy and irregular slip traces on the surface. The recent reviews and experiments of Vogel and Brick and Chen and Mad-din3 show that, for the body-centered-cubic metals, the resolved shear stresses along the planes and the degree of close-packing do not necessarily determine the slip system. Theirs and other investigations indicate that a complete understanding of the slip process in the body-centered-cubic metals is not yet possible. However, one behavior displayed by all these metals is that the slip direction is invariably one of the close-packed directions <11l>. The wavy nature of the slip-band traces has been explained on the basis of cooperative slip by planes sharing this same slip direction.1&apos;3l&apos; It is generally acknowledged that, when the hexagonal-close-packed metal magnesium (axial ratio c/a = 1.624) deforms by slip at room temperature, it does so by basal slip in the close-packed direction <1I%>. This type of slip has the conventional appearance of straight lines and is governed by the critical shear-stress law. Schmid and coworkers" showed that basal slip was operative in the temperature range of —185" to 300°C (—300" to 572°F). Work by Servi, Norton, and Grant6 has shown that, during the creep of coarse-grained aluminum at high temperatures, new slip systems come into operation. The existence of a new high temperature slip system at temperatures greater than 225 °C (437°F) for magnesium was suggested also by Schmid. This was later indicated by Bakarian and Mathewson&apos; to be slip along the pyramidal planes (10i1) in the close-packed direction <11%>.* They * Mention will be made in the text of pyramidal planes and prismatic planes. The pyramidal planes referred to are the type I. order 1 of the form {10il); and the prismatic planes of the type 1 of the form (10i0). For the sake of brevity, they are termed in the teat as pyramidal (10il: and prismatic (10i0) planes, respectively. Where reference is made to the pyramidal plane type 1, order 2 of the form {10i2), they are referred to as the pyramidal planes of the type (10i2:. observed that this kind of slip resulted in irregular markings on the specimen surface. According to them, a possible cause for the appearance of the markings was due to "limited accessory slip on the pyramidal {10i2) planes." Burke and Hibbard," on deforming a single crystal of magnesium at room temperature, found evidence of slip on a pyramidal plane {1011). They explained it as being due to the effect of grip restraints. It is pertinent to note at this point that, although basal slip has received a fair amount of attention, only two specimens have been investigated, that of Bakarian and Mathewson and of Burke and Hibbard, to establish the elements of nonbasal slip in magnesium. During the study of the deformation mechanisms

Jan 1, 1956

Through the courtesy of Mr. J. H. Batcheller, we are able to give the following extracts from a letter written by Mr. Caetani: It has been a nice idea of yours to collect for me all those scraps of news I received a few days ago and to send me such a long interesting letter. In four years war I have had practically no news. Burch wrote several times but several of his letters were lost and so mine. I believe I even wrote to you. Much has changed since our last days in Telluride; here and there many have disappeared. Of myself I will tell you in a few words: for a year I was in Cadore(?), where I had the good luck to do to the others what they wished to do to me, and had good luck on my side. Then I went to Macedonia for 5 months, did little fighting but lost my faithful fellow officer and companion.

Jan 6, 1919

By F. W. Bryan, D. H. Lynch

Introduction This paper is a general description of current central Florida phosphate mining, beneficiation, and product transportation. It is directed and believed to be of interest to engineers not familiar with this industry. Deposit: The phosphate deposits of central Florida are generally located in a five county area which includes Polk, Hillsborough, Hardee, Manatee, and DeSota counties. Geologically, the deposit is of marine origin and is identified as the Bone Valley formation. This formation is Pliocene to Recent in geological age and overlies a Miocene limestone formation known as the Hawthorn. The Bone Valley formation sediments are regionally characterized by equal proportions of apatite, quartz, and clay. The clay is predominantly of the mont-morillonite family. On a local scale, however, the proportions of these three major constituents vary considerably. The phosphate occurs as the apatite mineral (Ca 10F2(PO4)(6) and with the clay and sand, the minable ore is commonly referred to as matrix. This matrix is overlain by unconsolidated overburden of sand and sandy clays, ranging in depth from 10 to 45 ft. The matrix usually occurs in fairly horizontal continuous beds from 3 to 25 ft in thickness. The bedded limestone formation lies directly below the matrix and is generally well defined. The phosphate particles range from 3/4 in. to 200 mesh (Tyler) in size. The phosphate particles coarser than 14 mesh are called pebble phosphate and those less than 14 mesh are termed flotation feed which, when beneficiated, subsequently become concentrates. Through mining and beneficiation, phosphate quality is measured in BPL percent which stands for bone phosphate of lime units. In subsequent chemical manufacturing, the quality is indicated by P205 content. The deposit is economically characterized by various ratios such as tons of product per acre and cubic yards handled per ton of product. Magnesium, iron, and aluminum content are also considered in evaluating ore reserves. These elements are often critical to the chemical fertilizer processes. Presently, an ore body is considered economically minable if it meets the criteria shown in [Table 1]. These, of course, are general guidelines and specific costs and returns on investment must be considered in each case for acquiring reserves. On a new grass-root venture, a 20-30 year life is generally expected with a mineral recovery of 80%. History and Uses Phosphate mining in central Florida began around the turn of the century. However, in the early days, only pebble phosphate was produced until about 1930 when technology was available to beneficiate the -14 + 150 mesh particles. The -150 or -200 mesh material was discarded as it is today. The basic processes for beneficiation are washing, scrubbing, desliming, sizing, and flotation. These basic unit processes are essentially the same today although many improvements have been developed since the early days. Phosphate is used primarily in the production of high analysis fertilizer chemicals, typical of which are triple superphosphate, monoammonium phosphate (MAP), and diammonium phosphate (DAP). Phosphate is also used in the production of food preservatives, dyes for cloths, vitamin and mineral capsules, steel hardeners, gasoline and oil additives, toothpaste, shaving creams and soaps, bone china dishes, plastics, optical glass, photographic films, light filaments, water softeners, insecticides, soft drinks, road fill, and livestock feed supplements. Florida produces over 80% of the nation&apos;s marketable phosphate rock and one-third of the world production, according to the US Bureau of Mines. This amounted to approximately 35 million tons in 1975. Exports of Florida phosphate rock were to such countries as Canada, Japan, West Germany, Italy, and India, with Canada and Japan being the major users. Almost 95 o of all outbound cargo shipped through the port of Tampa is phosphate rock or related products. Beneficiation Following is a description of Agrico&apos;s new Fort Green beneficiation plant which is typical of the newer large capacity plants being built in the field. Agrico&apos;s Fort Green mine was completed in 1975 and is located in the southwest corner of Polk County and is directly adjacent to Manatee, Hillsborough, and Hardee Counties. With some minor differences, Fort Green is typical of a modern central Florida plant. The rated capacity is 3,000,000 plus tons of product per year and this varies according to the richness of the ore being handled. A simplified flowsheet is presented in [Figs.1 and 2]. This plant is served by three draglines of the 40-cu-yd class. The phosphate beneficiation is usually divided into three major functional steps: (1) washing and screening to produce a pebble product and flotation feed, (2) feed preparation and (3) flotation to produce concentrates. The typical plant is similarly divided into these three functional areas. Washer: Briefly, the slurried matrix is pumped from two draglines simultaneously at a combined rate of about 20,000 gpm at 2000 tph (solids) to rotary trommel screens sized to make a 7/8-in. separation. ([See Fig. 1]-) The trommel oversize is sent to hammermills where it is crushed and returned to the trommel screens, or pumped to tailings if minor impurities (Fe203, A1203, MgO) are too high. The trommel undersize is pumped to 14 mesh stationary (static) flat screens. The flat screen over¬size is subjected to three stages of 14 mesh vibrating screening and two stages of log washing in order to produce a final pebble product. The pebble product (+ 14 mesh material) is conveyed by belt conveyor to a large on-ground storage pile. Pebble product is reclaimed through a tunnel and loading system below

Jan 1, 1980

By E. A. Holbrook

The art of ore dressing is as old as civilization itself. Jason&apos;s search for the golden fleece was perhaps only the use of sheepskins for catching gold from gold-bearing sands washed over them. From Hoover&apos;s translation of Agricola we gather that in the middle ages ore dressing had advanced through the stage of hand sorting to water concentration by jigging. German methods and apparatus continued in the lead until the closing years of the past century; indeed, it is little more than a generation ago that many ore-dressing plants in this country were designed and operated by German-trained ore-dressing specialists. Meanwhile the development of the ore deposits in this country during the nineteenth century brought forth new inventions and methods. Blake&apos;s invention of the rock breaker in 1858 and the coming of the Wilfley table about 1896 are random examples of the improvements caused by our expanding needs. Yet ore dressing remained an art developed and practiced more or less by cut and try methods. The publication of the volumes on Ore Dressing by Dr. Robert H. Richards in 1903 marks the beginning of American ore-dressing practice as a science. Previously we had as guides only a few incomplete books based on foreign practice, or scattered references and articles in our own Transactions. Dr. Richards set down all the developments that had gone before, and outlined the ways for a scientific approach to the subject. He gave us a starting point. Since then ore-dressing practice and technique in this country have set the pace for the world. Coal preparation has been called "ore dressing upside down,&apos;&apos; because in ore dressing we generally save a small percentage of the heavy valuable mineral and discard the lighter gangue, whereas in coal preparation we save the larger quantity of light mineral and discard the smaller percentage of heavy waste. By comparison with ore dressing the history of coal preparation is brief, extending backwards hardly 100 years, for although coal had been dug and used locally for centuries it became an important article of commerce only with the coming of the industrial revolution 150 years ago. In new mines it was to he expected that the thickest and purest seams should be worked first. So long as these lasted there was little thought of coal preparation other than hand picking of the larger impurities and simple sizing to supply the real or fancied need

Jan 1, 1934

By Ashok K. Kakar

THE occurrence of superplasticity in Al-Zn alloys is of interest because of its possible industrial exploitation.&apos; The eutectoid composition (78Zn-22A1) has been the most extensively studied;&apos;-= neck resistant flow and high strain rate sensitivity of flow stress (characteristic of superplasticity) are associated with a micron-size grain conferred on the alloy by an homo-genization and quench.5&apos;6 Compositions away from the eutectoid have received comparatively little attention, especially from the viewpoint of relating the structural

Jan 1, 1969

By Henry Rogatz

During the year 1937, there were 663 oil wells drilled in the Texas Panhandle, making a daily initial production of 322,332 bb1.—that is, 191 more oil wells drilled than in the previous year, with an increase in the total daily initial production of 66,283 bbl. The total daily potential of the field as determined by the Texas Railroad Commission in its Schedule of Dec. 17, 1937, was 1,039,982 bbl., or an increase of 205,094 bbl. over the previous year. On Dec. 17, 1937, the daily allowable of 68,578 bbl. was assigned to the entire field. The total amount

Jan 1, 1938

MOWRY BATES, ? Tulsa, Okla.-In the first part of this paper the author says: "In an unpublished paper on Water Problem in the Augusta Field, S. K. Clark reaches the following conclusions: (1) That the great amount of water present is bottom water, occurring in the Varner sand, the main producing or the 2500-ft. horizon. (2) That the only striking connection between structure and water is in the area of the marked

Jan 4, 1919

By M. H. Harrington

The Transaclions of the Institute afford abundant evidence of the general recognition by mining engineers of the importance of a safety-lamp which will not only give warning of the presence of fire-damp, but will also indicate the proportion in which it is present.&apos; Mining-men are only too familiar with the frequent and sudden changes that habitually occur at the working-face of a gaseous mine. Nothing illustrates more forcibly the pregnant condition of the strata enfolding the seam and the seam itself, than the fact that a circulation of perhaps 150,000 cu. ft. of air

Jan 1, 1907

By Victor V. Damiano

THOMPSON1 proposed a very convenient nomenclature in which an imaginary regular tetrahedron was used to facilitate the analysis of dislocation interactions in fcc materials. The faces of the tetrahedron were constructed parallel to four different sets of (111) planes, and the six edges of the tetrahedron then represented the six 1/2 [110] Burgers vectors. The lines which joined the midpoint of a tetrahedron face with the ends of one of its edges represented the Burgers vectors of the partial dislocations into which the six 1/2 [I101 Burgers vectors dissociate. Using Frank&apos;s2 notation, all of the dislocations and their partial dislocations were identified and numerous possible interactions were postulated. The case of hexagonal crystals is unfortunately more complex, but it appears convenient to introduce a system similar to that of Thompson1 to describe the various slip planes and Burgers vectors encountered in cph structures. The system proposed has limited application but serves to identify the following slip systems {0001} <l20>, {1010} <1120>,

Jan 1, 1963

By T. M. Olson, B. G. Lewis

The adsorption of selenite on bentonite, kaolinite, and an iron--coated illite was studied to determine the importance of several soil variables on selenite removal from uranium mill tailings pond seepage. Adsorption capacities of the clays of selenite varied by orders of magnitude and decreased in the following order: illite > kaolinite > bentonite. Maximum adsorption of selenite on the clays was observed between pH 2.5 and the clay&apos;s zero-point-of-charge. Substantially less selenite was adsorbed on these clays in actual mill tailings solutions than in model systems. Virtually all selenite was removed from the tailings at pH 7. The principal removal mechanism at this pH is suggested to be the adsorption of selenite on freshly precipitated metal hydroxides. It is concluded that selenite can be removed from tailings seepage in soil or in clay liners in the presence of kaolinitic clays and iron oxides. Selenite would also tend to be removed from tailings seepage in calcareous strata.

Jan 1, 1984

By A. D. Hill, R. S. Schechter, M. P. Walsh, I. H. Silberberg, M. J. Humenick

There are many uranium deposits which are too deep, too poor in quality or too small in extent to be economically strip mined. It now appears that this uranium can be mined using in situ leaching processes. A number of field tests have been, are being or will be conducted to demonstrate the feasibility of In site leaching. Preliminary results are encouraging, but a major question concerns the restoration of the mine site, which is generally contained within an aquifer. It is necessary to ensure that the quality of the water is not adversely affected by the mining operation. One component of the lixiviant is necessarily a cation and, during the course of mining, the ion exchange sites of the minerals in the mine zone become saturated with this cation. Ammonium is the cation now commonly used. Clearly, any restoration procedure or question involving the rate of cation migration with the groundwater all necessarily involve the process of ion exchange coupled with flow in porous media. The equations describing this process are presented in this paper and numerical solutions obtained. Preliminary results showing the effect of isotherm shape, cation exchange capacity, flow and dispersion are presented. These results demonstrate that in many situations cation migration in ground water will be slow.

Jan 1, 1979

By J. H. Johnson, H. J. Schock

The cases of no auxiliary ventilation and ventilation with tubing were mathematically modeled, using a turbulent dispersion model. The Scooptram® used to load is modeled as a variable-speed, variable-strength source of pollutants moving in and out of the dead-ended drift. Model predictions are compared to experimental measurements. A parametric analysis was performed to determine If a relationship existed between C02 and NO levels. It was concluded from this analysis that when modeling a load-haul-dump cycle, it is necessary to use a model that can predict changes in source strength as well as source location. Carbon dioxide was used as a tracer to calibrate the model since, knowing the fuel flow rate, it is possible to calculate C02 production. The model was then verified after this calibration process using additional experimental results. With the calibrated model, it is possible to predict pollutant concentrations at any location in the heading using an engine map and the previously calculated flow field.

Jan 1, 1984

By W. H. Emmons

ALTHOUGH gold is one of the rarer metals, it is widely distributed; it is found on all of the continents and in each of the grand metallogenic provinces of the earth. It is prominent particularly in the shield areas of the earth; these areas have produced a large part of the world&apos;s lode gold and in most of them gold lodes are more important than the lodes of other metals. The Scandinavian shield is the only one that has not produced much gold. Nearly all of the world&apos;s gold has been derived from lodes and from sands and gravels where gold placers formed by the destruction of lodes. The gold derived from magmatic segregations, from pegmatites and from contact metamorphic deposits is quite subordinate to that derived from auriferous lodes. The lodes include all veins and related deposits that are formed by hot waters that move along fissures or other openings and that deposit vein matter in the fissures and by replacement of the country rocks along the fissures. All of the shield areas of the earth are characterized by large areas of granitic rocks including granitic gneisses, granodiorites and other acid intrusives which here are included in the general term "granite." The granites are intruded into schists, ancient lavas and other rocks which here are designated as the "invaded" rocks. The granites are parts of what once were great batholiths: that is, they are the remaining portions of great deep-seated bodies that slope outward, have very irregular roofs and so far as known have no floors but extend downward to great though unknown depths. Some of them are scores or hundreds of miles across and contain islands or "roof pendants" of the invaded rocks at places where the roofs of the batholiths were low-that is, where erosion has not yet completely unroofed the batholiths.

Jan 1, 1932

By M. W. Fabjanczyk, E. G. Alexander

INTRODUCTION The purpose of this paper is to summarise development of design processes used in the 1100 oreoody as a result of experience encountered. Design methods have been developed as a result of the history of ground behaviour, free-standing qualities of cemented back fill and mass blasting techniques. This paper contains an historical review of design methods previously described by Matthews ( 1972 ) , Hornsby and Sullivan (1977 and 19781, Kirkby, Blair and Hutton (19781, and Stewart (1980). The reasoning behind design philosphy changes are discussed and placed into context with methods now adopted. The 1100 orebody is Mount Isa&apos;s main producer of copper ore (5 million tonnes per year ) . Proven reserves are 120 million tonnes at 3.2% Cu. Location of the 1100 orebody in relation to overall mine geometry was well described by Dayton (1976). Geology The orebody is located in a series of massive brecciated shale lenses, in either a siliceous or dolomitic matrix, surrounded and separated by a uniform shale sequence. The whole formation is truncated at depth by a faulted contact with a greenschist facies metamorphic sequence below (Figure 1 and Figure 2) Mathias and Clark (1976). The breccia is known locally as silica dolomite and contains nearly all the economic mineralisation. Interfingering of shale and breccia often results in irregular boundaries to the silica dolomite and the ore limits. Geometry The geometry of the orebodies is shown by Figures 1, 2 and 3. Figure 3 also shows the location of the copper orebodies in relation to the lead orebodies. Structure Although the orebody is contained within the massive breccia (u.c.s. 150-250 MPa), it is cut by shear zones sub-parallel to the regional bedding (north-south striking, dipping approximately 60o to the west), and diagonally by major transverse faults (Figure A). These major faults have continuity in excess of 300 metres, with varying amounts of displacement up to 200 metres, and are zones of potential movement.

Jan 1, 1981

By D. L. Blakeman, Trimble. J. W., S. W. Banning

Noranda&apos;s Grey Eagle mine and mill, in the Siskiyou Mountains of northern California, began shakedown operations in the late fall and early winter of 1982. This paper describes some of the unique features in mineralogy, milling, and water treatment operation at Grey Eagle. In addition, major start-up problems in the mill and water treatment plant are discussed along with measures taken to correct these problems.

Jan 1, 1986

By P. K. Foster, A. McNabb

A consideration of the literature shows the inadequacy of Fick&apos;s Laws to describe the diffusion of hydrogen in steels below 400°C. These laws are modified an the assumption that hydrogen is delayed at fixed sites in the lattice. Approximate and exact solutions of the resulting nonlinear system are derived for simple geometrical shapes. The results qualitatively account for the observed anomalous behavior. Experimental methods for verifying the theory are suggested. 1 HIS paper presents a new analysis of the diffusion behavior of hydrogen in steel based on a physical model which supposes that hydrogen atoms wander in a random manner through the crystal lattice, but tend to get trapped or delayed at certain fixed sites uniformly distributed throughout the metal. These sites, which will be called traps, are regarded merely as potential wells of significantly greater depth than those encountered in regular regions of the crystal lattices and no assumptions are made concerning their nature or origin. The results of experimental studies of the diffusion of hydrogen in iron and steel have usually been reported in the literature in terms of a diffusion constant D derived under the assumptions that the hydrogen concentration C is governed by Fick&apos;s laws.1 Fig. I is a composite of published graphs of the man. ner in which the values of D, so derived, vary with temperature. It is convenient to consider this material in two categories; the high temperature results above 400°C, and the results at lower temperatures. The graphs 2, 3, 4, 5 in the first class lie within a factor of three of each other and are relatively consistent compared with the low-temperature values. There is an increasing divergence in values of D with decreasing temperature till at room temperature they are spread over three orders of magnitude. Small values of D have been obtained for work-hardened specimens.&apos; It is evident that D is a function of other variables besides temperature in this range and that these neglected variables are in some way connected with the work hardening experienced by the specimen. In view of evidence for an interaction between hydrogen in solution and dislocations9&apos;10 and adsorption on crack surfaces,&apos; some doubts must be entertained about the validity of Fick&apos;s laws and their associated physical model of noninteracting particles in random motion through the medium. More direct evidence of their inapplicability to the diffusion of hydrogen in steel was first produced by Darken and smith." Consider the permeation of hydrogen through a plate of thickness a under a boundary concentration Co at the ingoing, and zero at the outgoing, face. If Fick&apos;s first law1 is valid, the steady state flux P, through unit area is DCo/a. In addition, the concentration C varies linearly throughout and Qo, the mean quantity of hydrogen in the medium, is co/2. Compare Qo with Q1 the mean quantity required to saturate the same specimen to a uniform concentration G. In this case Ql = Co and hence the ratio

Jan 1, 1963

The Lornex mine is located in the Highland Valley area of British Columbia, Canada, one of the most,promising districts of low-grade porphyry-type copper mineralization in the world; Lornex is operated by Rio Algom Mines, Limited, a member of the Rio Tinto Zinc Group of London. Exploration from 1965 to 1968 disclosed an ore body of 293 million tons with a grade of 0.43% copper and 0. 14% molybdenum; this led to the decision to construct a 38,000 tpd operation, and production commenced in the fall of 1972. The ore is open-pit mined with 15 cu yd (11 5 m3 shovels, which load 100- and 200 -ton trucks. These trucks dump directly into a 60 in. by 89 in. Allis -Chalmers gyratory from two sides of the 14 ft 6 in. (4.42 m) deep dump pocket and across splitters mounted on the pocket sides. The unlined pocket would hold 1,000 tons level full. The crushedore discharges into a 300-ton surge hopper equipped with an automatic sonic level indicator to prevent over -filling. A 96 in. by 23 ft (2.44 m by 7 m) apron feeder delivers the ore from the surge hopper to a 72 in. (1.83 m) transfer belt equipped with a tramp iron metal detector. The transfer belt discharges to a 60 in. (1. 52 m) conveyor which conveys the ore about 3,000 ft (914 m) to the 475,000 ton ore storage pile located near the concentrator. Four 48 in. (1.22 m) pan feeders transfer the ore to the mill feed conveyors, which in turn supply the two 32 ft by 15 ft (9. 75 m by 4. 57 m) semi-autogenous mills, which easily accept the 10 in. (254 mm) product of the crusher. Each semi- autogenous mill is followed by a secondary ball mill. The unusually large open side setting of the crusher accounts for its 6,300 tph maximum capacity. Concentrator capacity is now 48,000 tpd. Eccentric removal is eased by the design of the door to the crushed ore pocket, which has the cart rails mounted on its outer surface and is lowered to a horizontal position in the pocket by an electric winch. The top of the door rests on wear bar covered concrete piers in the pocket. Another feature is a chute through the floor at the crusher base that extends into the crushed ore pocket. Dust and spills of ore on that floor during crusher repairs are easily swept into the crushed ore pocket. This feature is also used at Bougainville and Palabora and has been added to others after operation has commenced. The design follows the concrete rectangular box concept with four main service and maintenance levels shown in the sectional drawings, although a third additional section has been added at the maintenance

Jan 1, 1978