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By Ta M. Li

Application of an hydraulic mining system to a troublesome overburden removal project has unlocked an estimated three million tons of ore reserves at Caland Ore Co., Ltd.'s open-pit iron operation in Ontario, Canada. The mine, operated by Caland under a 99-year lease and royalty agreement with Steep Rock Iron Mines, Ltd., would have been forced to curtail its mining operations later this year as the economic limits of the original pit design were being reached. As a result of the stripping program, which uses a tandem set of Marconaflo jet slurry systems, Caland officials now estimate that mining will continue into 1977. These reserves, added to existing ore stockpiles, are expected to sustain the firm's pelletizing operation well into 1981. In addition, Caland engineers are initiating studies to continue mining since the ore body, to date, has yet to bottom out.

Jan 10, 1976

During the year 1915-16, the Mining Engineering Society of the, Massachusetts Institute of, Technology had five meetings, at which the following gentlemen addressed the society on the subjects Mentioned: Professor Richards-Memory and the Business World. Mr. Hutchinson-Mine Examinations. Mr. Packard-Mining Law, Necessity of Amendment. Mr. A. H. Rogers-Organization of Mining. Property. Mr. C. Hurter-Explosives. Mr. W. Macgregor-Organization of a Gold Mine and Mill. During the year 24 new members were admitted, making the present active membership 54. There was an average attendance of 34 members at the meetings. HAIG N. SOLAKIAN, Secretary, 1915-16. The Pick and Shovel Club, Case School of Applied Science, has elected the following officers for the current year: President, G: D. Welty, Secretary, C. E. David, Vice-President, E. G. Hollman, Historian, L. S. Kuehn, Treasurer, C. E. Althouse, Sergeant at-Arms, H. H. Smith.

Jan 2, 1917

Three years ago, when Sun Energy Development Co. (Sunedo) dedicated its Cordero mine in Wyoming, the company noted that its decision to venture into the coal industry was prompted by the need for domestic energy self-sufficiency. Now, as the energy situation assumes crisis proportions and the cry For coal intensifies, Sunedoo has begun a $21 million expansion program that will double the mine's productive capacity by the mid-1980s. When the Cordero come was originally conceived, Sunedoo obtained permission from permitting authorities to produce up to 11 million t/y (12 million stpy) of coal. But seeing opportunities in the coal market, Sunedoo returned to the agencies in 1976 and came away with permission to produce 22 million t/y (24 million stpy). Of that total permitted capacity, Cordero is expected to produce about 3.6 million tons (4 million st) this year.

Jan 9, 1979

By Paul T. Allsman

Recently the U.S. Bureau of Mines completed organizational changes that, for the first time since the beginning of World War 11,. re-emphasize mining research as an important part of its responsibilities. A progress report follows on the Bureau's improved program for mining research. During World War II the Bureau's mining research lapsed in favor of short-range, more urgent activities. Since the end of the war mining has grad- ually returned to an economy where reduced costs and increased efficiency are essential to a vigorous industry, and demand for improvement through expanded research is again pre-eminent. In recognizing this situation, the Bureau is again emphasizing mining technology and has reinstituted some of its earlier mining research activities. Recent research, however, follows a more scientific approach in solving mining problems because new technology, particularly new instrumentation, has developed rapidly during the past decade.

Jan 7, 1960

By Robert B. Schlosser

Anaconda's Jackpile uranium mine is a multi- million-ton open pit sandstone deposit, producing some 3000 tpd for the company mill. With minor exceptions, the orebody is in equilibrium, and the backbone of ore-grade control is the radiometric assay, which fortunately is dependable. Preparatory Sample Drilling: Before it is stripped, the orebody is diamond-drilled on 200-ft centers N-S and 100-ft centers E-W. Drill core is logged, split, and assayed radiometrically, and several samples are assayed chemically to check the equilibrium status. This information establishes the pit outline and the first rough estimate of ore. After stripping, diamond drill data is supplemented by rotary drillholes on 50-ft centers. Dry chip samples, collected every 2 ft, are prepared and assayed radiometrically at the mine. Details of the orebody are then available for study, and the basic mining plan is drawn up.

Jan 10, 1958

Members of the American Institute of Mining and Metallurgica Engineers can now secure copies of the Journal of the Institute of Metals of Great Britain at a special price of $5 for the two volumes issued during the year, postage extra amounting to approximately 25 cents. This Journal is one of the best books dealing with the subject of non-ferrous metals and alloys and this arrangement with the Institute of Metals permits our members to take advantage of this specially low price. Back numbers of the Journal can be purchased on the same basis. The Journal can be ordered directly through the office of the Institute of Metals, Great Britain 36 Victoria St., Westminster, S. W. I., London, England. The letter should be accompanied by an international money order and a copy sent to Secretary of the A. I. M. E. in order that the membership record may be checked.

Jan 9, 1919

Department of Forestry and Geological Development, 435 State Capitol, Atlanta, Ga. S W. McCalhe, State Geologist. A complete list of Bulletins of the Geological Survey of Georgia will be sent upon request. Forty-five Bulletins have been issued, most of which deal with the mineral deposits of the state and are still available; they may be had on payment of postage. The subject matter of those bulletins of most interest is indicated in the following list, with the year of issue and the mailing weight. Bulletin 1, Marbles (1894), 3 lb , 8, Road building materials (1901), 3 lb , 9, Granites and gneisses (1902), 2 lb ; 10, Iron ores (1900), 2 lb , 11, Bauxite (1904), 2 lb.; 12, Coal deposits (1904), 2 lb ; 13, Ocher deposits (1906), 2 lb.; 14, Manganese deposits (1908), 2 lb ; 17, Fossil iron ore (1908), 2 lb.; 18, Clay deposits

Jan 1, 1933

By W. C. Hagel

In 1952, Hauffe and pfeiffer1,2 reported that parabolic rates of nickel oxidation are decreased by factors of two and four at 1100° and 1000°C, respectively, when Li2O vapor is present. Hauffe 's3 explanation was that monovalent lithium enters the crystal lattice of p-type NiO to decrease cation-vacancy concentration and, hence, diffusion of Ni2+ ions across the growing oxide. After first succeeding in adding lithium directly to molten nickel, the purpose of this investigation was to determine whether alloyed lithium is a more effective oxidation inhibitor than the external vapor. Seven 650-g heats were prepared by melting carbonyl nickel inductively under 1 atm of argon. Various amounts of lithium were added by dropping small chunks into the melt. Only about 10-pct recovery was achieved with the remainder lost by spattering and volatilization. Nominal and actual

Jan 1, 1965

By John W. Sweeney

Florida is blessed with some of the worlds greatest phosphate rock deposits, and doubly blessed in that those deposits are uraniferous. Until recently, the uranium con¬tained in phosphate rock had been ignored, but spot mar¬ket prices in the $40-per-pound range have spurred actions which will soon transform the state into a major uranium supplier. In 1978, two plants in Florida and one in Louisiana went it on- stream to recover uranium oxide from phosphoric acid, and construction began on three other extraction facilities. Together, these six plants will be cap¬able of recovering 1940 t/y (2140 stpy) of yellow cake from Florida phosphate in the early 1980s, equivalent to 15% of the projected US requirement. By the year 2000, Florida phosphates will have produced a projected 49 000 tons (54,000 st) of uranium oxide to fuel the nation's nuclear reactors.

Jan 9, 1979

By Harry M. Mikami, A. Gene Sidler

Chemistry of the evolution of materials in contact with copper converter tuyeres is delineated by means of analyses of periodic punch rod samples taken during a converter cycle. Lining samples from known peripheral and end locations from three converters were investigated chemically and petrographically. One set of eight specimens representative of all lining parts was removed for study before completion of the campaign. Purely surface removal by slag action is minor. Slag filling of interconnecting cracks extending to the hot face results in a "stoping'' action which shapes the hot face contour and is a significant factor. Major wear mechanism is believed to be copper metal slabbbing caused by metal first penetrating joins and invading fractures or impregnating brick structure laterally from the joints. Importance of lining design and joint fit are emphasized. New laboratory tests are indicated. In the copper converter, molten matte, which is essentially a solution of iron and copper sulfides, is blown with air to produce blister copper. During the first stage of the process the FeS component is oxidized to SOz and FeO, the latter combining with a siliceous flux to form a predominantly ferrous silicate slag. Successive increments of matte and flux are charged, blown, and skimmed of slag until sufficient CuzS ("white metal") has accumulated. This white metal is then blown to copper metal in the final or blister-forming stage by the oxidation of sulfur to SOz and the reduction of combined copper to metal. The reactions of both stages are exothermic so provide the required heat. Temperature is controlled by cold scrap copper, flue dusts, smelter reverts, and so forth. Considerable vari- ation in practices occur depending on the type of converter, grades of the matte and flu, and other factors. Lathe and Hodnett's' survey of world converter practices gives much information on these variations and their effects on the refractories. Principles of converting chemistry are presented by Newton and wilson,' Ruddle,' and schuhmann. Most copper converting occurs within the temperature range 2100" to 2370°F although individual smelters report1 ranges from as low as 1700" to 2240°F to a high of 2200" to 2400°F. Temperature intervals reported from beginning to end of blow varied from 45" to 540°F. Average temperature range was 2030" to 2260°F and the average interval 250°F During the matte-blowing stage varying amounts of magnetite (FeO.Fe,O,) are formed in the slag as suspended crystals and also the liquid portion of the slag will contain, in addition to ferrous silicate, some Fe20s which adds to the potential magnetite formation if the temperature falls. This is a component of the so-called "dissolved magnetite" often referred to in the literature. Although the slag may contain minor amounts of CaO, A1209, MgO, and alkalies, the principal components are Fe0-Fez03-SiO,. The phase equilibria relationships of this system have been elucidated by Muan~ and Schuhmann, et a1.' These data and estimates of converting atmospheres by schuhmann4 are useful tools in studying slag behavior. It would appear that in many cases reported magnetite contents of slags are apparently higher than actually is present at operating temperature. This is probably because much of it is formed during slag solidification. The quaternary systems which have Mg-0,1,0,' and addition to iron oxides and silica are also of interest in considering the reactions of the slags with basic refractories. The formation of magnetite has an important interrelationship throughout the pyrometallurgy of copper (Mossman,' Gronningsater and Drummond," Ruddle,' Elwood and Henderson") but its presence in the converter particularly relates to operating efficiency and lining life in certain instances. Archi-

Jan 1, 1963

By N. Mungan

A study was made of the effects of wettability and interfacial tension on the immiscible displacement of a Iiquid by another Iiquid from porous media. The influence of viscosity ratio was also investigated. Porous media used were polytetrafluoroethylene (TFE) cores prepared by compressing TFE powder under different pressures. lt is shown that displacement of a wetting by a nonwetting Iiquid is always less efficient than the displacement of a nonwetting by a wetting fluid, all other things being equal. In the former case, the recovery efficiency can be increased substantially by either reducing the interfacial tension or increasing the viscosity of the displacing fluid. A qualitative discussion is given on the implications of this work to the recovery of crude oil by water flooding. INTRODUCTION The high cost of oil exploration and new recovery schemes makes it imperative that waterflooding be conducted under conditions favoring most efficient oil recovery. To improve oil recovery by waterflooding, it is essential that the role played by interfacial forces in the entrapment of residual oil be studied and understood. Interfacial phenomena in natural rock, connate water and crude oil systems are very complicated because of the complexity of the natural liquids found in petroleum reservoirs, because of our inability to adequately describe the geometrical structure of the porous media and because of a lack of understanding of physical and chemical interactions between the liquids and surface of the pores. The problem becomes further complicated when one tries to elucidate the role of interfacial phenomena in fluid flow. Numerous studies of the displacement of oil by water under different interfacial tension or wet-tability conditions have been made.lm3 These studies have been performed in silica, alundum or sandstone systems using water and paraffin oil and also some surface active material to control the interfacial tension s and the contact angle 0. Unfortunately, the high energies of various interfaces involved favor adsorption and orientation of the surface active material at the interfaces. Also the surface active material concentration at the interfaces exceeds that in the bulk of the liquid phases. Such surface excess may cause the surfactant distribution, the contact angle and the interfacial tension to differ from their measured static equilibrium values and makes interpretation of the displacement experiments difficult. Furthermore, as changes in s also lead to changes in cos 0, the role played individually by one of these parameters in the displacement becomes obscured by the effect of the other. To circumvent these difficulties, a low surface energy solid and true solutions or pure liquids should be used. Use of a low surface energy solid minimizes adsorption and orientation effects at the solid-liquid interfaces. By controlling s and cos 0 through use of selected pairs of pure liquids or true solutions rather than by surfactants, the adsorption effects at liquid-liquid interfaces are eliminated. In the present study TFE cores were used as the porous media. Liquids used were water, sucrose solutions, paraffin oils and benzyl, n-butyl and isobutyl alcohols. The interfacial tension was varied from 40 to 1.1 dynes/cm by suitably choosing the liquid pair. A surface active material was added to the water-oil system only in the case where interfacial tension of 0.5 dyne/ cm was desired. No precise changes of cos 0 were attempted. However, either the displaced or the displacing liquid could be made the one which preferentially wets the TFE surface. Using sucrose solutions and blends of paraffin oils proved to be a convenient way of changing the viscosity ratio between the displaced and displacing liquids. The present investigation examines the effect of interfacial tension, wettability and viscosity ratio on the immiscible liquid-liquid displacement from porous media. Observed data permit a

Jan 1, 1967

By Arthur H. Wilson

INTRODUCTION The stratified deposits of the coal measures are strongly laminated and exhibit a high degree of anisotropy. The stronger rocks in the sequence possess joints and shear planes, the coal itself has a closely spaced cleat pattern, and the softer rocks are capable of flowage. To calculate the general distribution of stress in such a medium by analytical means based on the theories of elasticity is difficult, and has not yet been satisfactorily achieved. Nevertheless, some estimate is required to aid the design of supporting pillars and permit the correct siting of main roadways. This is important when laying out districts for longwall extraction, particularly if the results of previous experience are not available in the area to be exploited. Although elastic stress analyses, including finite element analysis, may not be very appropriate, the basic equilibrium conditions will apply. These, combined with logical deductions based on observed fact, can be used to give appropriate solutions which experience has shown to be capable of giving acceptable results in the design of pillars associated with longwall workings. Constancy of cover load is one of the fundamental equilibrium conditions. When mining below ground, the cover is not removed, therefore the average vertical load after mining must equal the initial cover load. If mining produces a local decrease in load in any particular area, it must be balanced by an equivalent increase in load nearby. Bending moments in the strata above are capable of compensating for the turning moments which this redistribution of load may produce, and the existence of a "rear abutment", much discussed between 1950 and 1960, is not an essential equilibrium requirement. If the distribution of vertical stress is plotted three-dimensionally, then the load on any small element of area 6S will be o.6S, where o is the average stress. This also represents the volume of the "stress envelope" which lies above 6s. By making the individual areas very small and summing together, it is possible to show that the total load on any particular area can be found by calculating the volume of the stress distribution envelope which lies above it. A similar consideration applies to the stress distribution in a two-dimensional situation. If the stress distribution curve is drawn, then the area below any portion will represent the load per unit depth of section. For example, Figure 1 represents the possible stress distribution across a caved waste and associated ribside. If AW is the load deficiency [ ] These loads must balance, therefore area AW = AS, and if one can be deduced, the other will then be known. Attempts have been made to measure stress distributions in the ribside and in the waste. Stress meters in soft rocks such as the coal measures have not proved very effective. Those based on the measurement of strain have experienced difficulty in strata capable of flowage, high modulus stress meters suffer from cross sensitivity, and overcoming to obtain absolute stress is almost impossible. Load measuring instruments left in the waste have also proved largely ineffective, mainly as a result of the difficulties of protecting the leads in a caving environment. Certain logical deductions, however, can be made about the conditions in the waste. CONDITIONS IN THE WASTE Over the area of the waste the surface trough which develops is remarkably consistent, despite the geological variations which occur from coalfield to coalfield. In Britain, an accuracy of 210% is claimed for subsidence prediction based on width of extraction and depth of cover (National Coal Board, 1975). The subsidence in the USA is more variable (Adamek and Jeran, 1981), particularly with respect to the point of inflexion, but the same general pattern occurs. At the shallower depths the somewhat harder roof strata involved in the caving in the USA may suffer less compaction, giving reduced values of subsidence. Subsidence at intermediate horizons below the surface

Jan 1, 1982

By Peter A. Allen

INTRODUCTION Lac Minerals Ltd., a significant North American producer of gold and other metals for over three decades, recently completed the financing for its third discovery in an eight year period. From 1975 to 1982, three important gold deposits in Quebec and Ontario: namely Bousquet, Doyon and Hemlo, were located and tested by company geologists. When current development plans are complete, Lac will have injected $400 million into these deposits. A discussion of these projects and their financing reveals some extraordinary challenges faced by top mining management. Responses to these challenges developed by Lac management may interest executives planning to proceed with new North American gold or base metal mines. MINE DE BOUSQUET In 1976, we explored a property containing a large low grade surface gold deposit. The geologial setting was exciting and we tested recent rethinking about the volcanigenic genesis of gold and base metal occurrences. These ideas helped pus to persist in drilling a nearby conductor and 600,000 ton, 0.22 ounce/ton orebody was outlined. With gold at $Cdn.150 per ounce the feasibility was approved and work started February 1978. Availability of a company custom mill 35 miles away reduced our immediate capital needs to $8.5 million. First of all, understanding the geological significance of results at Bousquet was critical. Mine and surface exploration geologists at Lac are continually and closely working together to better understand the nature of gold deposits in the Shield. This management emphasis on both pure and mine geology meant that a much larger ore reserve could confidently be predicted. (A proposal to quickly develop and ship the ore to a nearby company mine was discarded.) Instead of designing just for the initial plan of 500 tons per day we therefore included headframe, shaft and hoisting capacity for 1,700 tons and provided a compartment for future shaft deepening. These prebuilt additions were welcomed within two years. The type of collaboration whereby surface and mine geologists make study and field projects of each other's problems sets the foundation for this predictive ability. At Lac, the traditional responsibility for and supervision of the mine geologist by the mine manager is modified by this collaboration. This adds interest and scope to the geologist's work and lets him think beyond production requirements. Besides such management considerations, dedicated, professional geologists with integrity and lots of practical experience are needed to plan successful developments like Bousquet. Second, projecting future bullion revenue was critical and our experience selling production for future delivery gave us a way to ensure that loan repayment at term would not be jeopardized by falling metal prices. Output from a single mine is often reduced or curtailed for numerous reasons, so "selling forward" is done best and with least risk by companies like Lac with more than one or two producing locations for that metal. At first, a forward marketing program was considered unwise, speculative and unorthodox by analysts and other producers. Declining markets in the early 1980s prolonged gold project capital recoveries however, and critics were better able to appreciate the usefulness of such a tool. The notion that gold miners shouldn't sell forward is rooted in the belief that long-run gold prices will rise as modern social theory implementation in democracies inevitably erodes currency values. Selling against this trend goes against this industry belief or "ethic" and is at odds with the company shareholder who expects the gold price to rise not fall. Nonetheless, an efficient tool is available to the gold producer (not to most other metal producers) to quickly disengage from adverse price trends or to lock in spot prices instantaneously for two years or more of profits. Gold had recently retreated from U.S.$200 to $104 and now had reversed to $150. Lac had been forced to close several of its producers during the late 1960s because central bank-controlled

Jan 1, 1985

By J. S. Browning

The U.S. Bureau of Mines has been experimenting with flotation processes to separate the spodumene-beryl ores mined at Kings Mountain, N.C. The success to date as well as the present status of the process is discussed. Further experimentation is underway and studies will be made on the milling costs involved in the flotation process. The pegmatites of the Kings Mountain-Lincoln ton, N.C., area constitute the largest known domestic reserve of beryl and spodumene. The reserve is estimated to contain 90 million tons of pegmatic material with 1,280,000 tons of recoverable Li,O(lithia) as spodurnene.' The pegmatites also contain 0.4 to 0.5 pct beryl disseminated throughout the orebodies. The pegmatites may contain a potential reserve of 240,-000 tons beryl, equivalent to approximately 34,000 tons of BeO. Different flotation methods for separating beryl from feldspar or quartz, or both, have been developed by various investigators.2-7 On the other hand, published information is limited on concentration of spodumene-beryl ores. As the response of beryl and spodumene to flotation is essentially the same, a successful separation of the two minerals depends upon use of a selective depressant or a selective collector for one of the minerals. Several years ago, the U.S. Bureau of Mines, recognizing the need for developing a large domestic supply of strategic beryl, undertook studies to develop improved methods of recovering this mineral. Because of the low beryl content of the pegmatites, economic production of the beryl entails its recovery as a byproduct of spodumene flotation. Thus, the development of procedures that would provide for maximum recoveries of the spodumene, as well as the beryl, was a primary objective of the studies. Tests were made on run-of-mine ore and spodumene flotation tailing from the Foote Mineral Co.'s spodumene concentrator at Kings Mountain (Fig. 1). Pet-rographic analyses of the ore and flotation tailing are given in Table I. Detailed analyses and examinations of the ore and tailings revealed that the beryllium was present in the form of a low-alkali beryl. The crystals were small, rarely larger than 14-mesh diam, and were clear and colorless. About 10 pct of the Li20 content of the ore was present in the mica and feldspar components of the pegmatite. In some weathered ores the associated clays also contained small amounts of lithia. This is typical of pegmatites in the Kings Mountain area. LABORATORY TESTS In the past, a vigorous chemical treatment of the spodumene surfaces with acid or caustic soda has been considered essential for satisfactory selective flotation of the spodumene from the other minerals in pegmatites. Many tests were made during this investigation to determine if chemical treatment of the spodumene could be eliminated in favor of a reagent combination that would depress the other minerals while selectively floating the spodumene. A simple reagent combination and process was developed that gave better spodumene recovery and grade of concentrate than the more elaborate acid or caustic treatment methods previously used. The finely ground ore pulps were conditioned with either an ammonium, alkali, or alkaline earth lignin sulfonate and sodium fluoride, and the spodumene was floated and cleaned using oleic acid as the collector. The spodumene tailing was deslimed, conditioned with sulfuric acid and coco amine acetate, and floated to reject mica. The mica tailing was thickened and conditioned with hydrofluoric acid, then washed to remove the acid. The washed pulp was conditioned with sodium hydroxide for pH control, and the beryl was floated and cleaned with oleic acid as the collector. Table II summarizes the results of the laboratory test work. About 77 pct of the spodumene and 75 pct of the beryl were recovered from the pegmatitic material. The spodumene concentrate assayed 6.1 pct Li20 and 0.01 pct BeO, and the beryl product assayed 3.0 pct Li20 and 1.57 pct BeO. Similar results were ob-

Jan 1, 1961

By G. R. Smithson, John E. Hanway

This paper, the first of two papers covering the development of a metallurgical process for the treatment of a complex ore, describes the bench-scale experimental Program undertaken as a basis for initial process design. The results of using various potentially applicable process steps are detailed and the basis for selecting certain of these steps for Pilot-plant development is outlined. A process consisting of the steps of sulfate roasting, leaching, copper recovery, solution purification, zinc electrolysis, and brine leaching for lead recovery was recommended for further study. THIS paper describes the bench-scale development of a sulfate roasting process designed to recover copper, lead, and zinc from a complex ore. The extrapolation of the bench-scale results to pilot-plant design and the results of pilot-plant operation will be described by Lund, et al.' in a subsequent paper. Several years ago, the research divisions of the St. Joseph Lead Co. became responsible for the development of a metallurgical procedure to treat a complex sulfide ore containing copper, zinc, and lead sulfides. The primary objective was to devise a procedure which would permit economic recovery of the nonferrous values. As could logically be expected, the study of conventional flotation procedures was undertaken; in addition, however, because the physical nature of the mineral association in the complex ore indicated that physical methods of ore beneficiation might not prove to be entirely satisfactory, a program also was established to investigate alternative methods of direct metallurgical treatment for recovery of the metal values. Exploratory test work by the St. Joseph Lead staff led to the conclusion that a sulfate roasting technique represented the most attractive approach for a direct metallurgical process. Sulfate roasting would provide for the selective conversion of the nonferrous values to forms soluble in conventiona1.1 leaching agents and would permit their efficient separation from the insoluble iron and other gangue materials. It should allow efficient utilization of the thermal energy available from oxidation of the sulfide raw material, and might provide an opportunity for recovering byproducts such as iron and sulfur. The decision to evaluate a selective sulfation procedure as a basic processing step also was influenced by the excellent results obtained by the St. Joseph Lead Co. in using this procedure for the recovery of cadmium from zinc sulfide concentrates. The choice of selective sulfation as a major processing step served as a basis for an extended experimental investigation of the process with bench- and pilot-scale equipment to prove its feasibility. St Joseph Lead Co., aware that Battelle Memorial 1nstitute had participated in the development of several fluidized-bed sulfate roasting procedures, asked Battelle to undertake the bench-scale development of a suitable sulfation process. This paper describes the procedure used and the results obtained at Battelle in the treatment of the complex sulfide ore. EXPERIMENTAL WORK Nature of the Raw Material. The sulfide ore deposits of interest are essentially a massive sulfide, with the indivudual ore minerals being extremely fine grained and intimately intergrown with each other and with the associated gangue. The zinc in the ore is present mainly as sphalerite (ZnS), the lead as galena (PbS), and the iron either as pyrite (Fes2) or as pyrrhotite (FeS). Table I shows the chemical analysis of various individual samples of the ore. These analyses illustrate the variability of the ore body and impose a rather severe requirement on any type of metallurgical process—that of being able to tolerate probable wide variations in the assay of the head ore. The tentative flowsheet designed to serve as a guide for the initial experimental investigation of

Jan 1, 1962

By B. Haimson, C. Fairhurst

A criterion is Proposed for the initiation of vertical hydraulic fracturing taking into consideration the three stress fields around the wellbore. These fields arise from (1) nonhydrostatic regional stresses in earth, (2) the difference between the fluid pressure in the wellbore and the formation fluid pressure and (3) the radial fluid flow through porous rock from the wellbore into the formation due to this pressure difference. The wellbore fluid pressure required to initiate a fracture (assuming elastic rock and a smooth wellbore wall) is a function of the porous elastic constants of the rock, the two unequal horizontal principal regional stresses, the tensile strength of the rock and the formation fluid pressure. A constant injection rate will extend the fracture to a point where equilibrium is reached and then, to keep the fracture open, the pressure required is a function of the porous elastic constants of the rock, the component of the regional stress normal to the plane of the fracture, the formation fluid pressure and the dimensions of the crack. The same expression may also be used to estimate the vertical fracture width, provided all other variables are known. The derived equations for the initiation and extension pressures in vertical fracturing may be employed to solve for the two horizontal, regional, principal stresses in the rock. INTRODUCTION Well stimulation by hydraulic fracturing is a common practice today in the petroleum industry. However, this stimulation process is not a guaranteed success; hence, the deep interest shown by the petroleum companies in better understanding the mechanism that brings about rock fracturing, fracture extension and productivity increase. Geologists and mining people became interested in hydraulic fracturing from a different point of view: the method may possibly be employed to determine the magnitude and direction of the principal stresses of great depth. Numerous articles in past years have dealt with the theory of hydraulic fracturing, but they all seem to underestimate the effect of stresses around the wellbore due to penetration of some of the injected fluid into the porous formation.1-4 Excellent papers on stresses in porous materials due to fluid flow have been published5,6 but no real attempt has been made to show the effect of these stresses in the form of a more complete criterion for vertical hydraulic fracturing initiation and extension. This paper is such an attempt. ASSUMPTIONS It is assumed that rock in the oil-bearing formation is elastic, porous, isotropic and homogeneous. The formation is under a nonhydrostatic state of regional stress with one of the principal regional stresses acting parallel to the vertical axis of the wellbore. This assumption is justified in areas where rock formations do not dip at steep angles and where the surface of the earth is relatively flat. This vertical principal regional stress equals the pressure of the overlying rock, i.e., S33 = —pD, where S33 is the total vertical principal stress (positive for tension), p is average density of the overlying material and D is the depth of the point where S33 is calculated. The wellbore wall in the formation is considered to be smooth and circular in cross-section. The fluid flow through the porous elastic rock obeys Darcy's law. The whole medium is looked upon as an infinitely long cylinder with its axis along the axis of the wellbore. The radius of the cylinder is also very large. Over the range of depth at which the oil-bearing formation occurs, it will be assumed that any horizontal cross-section of the cylinder is subjected to the same stress distribution, and likewise that it will deform in the same manner. STATE OF STRESS PRIOR TO DRILLING OF WELLBORE At a depth D from the earth surface within the oil-bearing strata, it is assumed that the following are 1References given at end of paper.

By E. J. Ripling, S. I. Liu

IT has been shown in a number of recent publications that much information on the mechanical behavior of metals can be gleaned by first deforming test specimens under one set of conditions, and then evaluating the effect of this deformation by means of a subsequent tensile test. A straining procedure of this type has been applied to study the flow behaviors of a number of metals.1, 2 In these studies the prestraining and testing were both tensile strains, while the straining temperature was varied between the prestraining and testing steps. These data indicate that there is a more or less general flow behavior exhibited by all metals which are pre-stretched at one temperature and then tested in tension at a second temperature. Prestraining a metal at a high temperature produces a lower yield strength during the subsequent low temperature strain than would the same total strain at the lower temperature. On reversing the order of the strains, low temperature prestraining produces a higher yield strength during the subsequent high temperature straining than would a single high temperature strain of the same magnitude. Unlike the flow behaviors, the fracturing characteristics of metals when prestrained under one set of conditions and tested under a second set of conditions is quite material dependent. The fracture behavior of a number of steels2" as well as of a commercially pure zinc,4 have been investigated by these two-step tests and was found to depend upon the position of the transition temperature with respect to the prestraining and testing conditions. Since these fracturing phenomena are complicated by the appearance of the transition temperature, it has been difficult to reliably define the basic effect of a two-step straining procedure in fracturing. Metals crystallizing in the face-centered cubic system do not exhibit transition temperature effects so that the mechanical characteristics obtained on these materials could serve as a base for two-step straining behaviors. Consequently groups of specimens of the aluminum alloy 24s-T4 were prestrained various amounts at room temperature in either tension or compression after which these strained specimens were further strained to failure, or tested in tension, at either room or some lower temperature. Other groups of 24S-T4 specimens were prestretched at various low temperatures after which these were tested in tension at room temperature. These straining techniques then made it possible to compare the behaviors of aluminum with the behaviors previously presented for steels. Material and Procedure The commercial aluminum alloy 24S-T4 in the form of 3/4 in. rods was used in this investigation. The material was subjected to the following treatments before machining: 1—resolution treating at 920" & 10°F for 45 min, 2—quenching in water at room temperature, and 3—aging at room temperature for four days. The re-heat treated material was then machined to either of the specimen shapes shown in Fig. 1. Fig. la shows the threaded-end specimen used for prestraining in compression. Fig. 1b shows the specimen used for both prestrain in tension and final testing. The material subjected to precompression in the form of the specimen shown in Fig. la was re-machined to the specimen shape in Fig. 1b for final testing. Prestraining in tension and final testing were done without any intermediate machining. All straining was performed on a 10,000 1b Riehle tensile testing machine. Tensile strains were conducted in a specially designed concentric fixture which produced an eccentricity of less than 0.001 in." The change in diameter at the minimum section of

Jan 1, 1954

By L. M. Yokum

Tremendous strides have been made in front end loaders in the last 15 years. Fifteen years ago a 21/4-cu-yd front end loader was the largest standard machine you could purchase. Today there are two commercially available 15-yd front end loaders and many 10-yd loaders. The estimated total worldwide market for all front end loaders over 1 cu yd has increased to over 30,000 units yearly. The trend for front end loaders, in general, and for the mining market, in particular, is to larger and larger loaders. Today there are 18 and 24-cu-yd front end loaders in prototype form. It is not inconceivable that within another 15 years there will be front end loaders in the 50-cu-yd class with design features quite different from today's. These larger loaders are being designed at the demand of the mining industry, and the design criteria are being directed at this market area. The costs of these larger loaders may be greater per unit of capacity, but it is the intent of the various manufacturers that they will be more reliable, be easier to service, give greater availability, live longer, and cost less to maintain, than previous loaders. It is also the intent of the manufacturers to compete directly with the large mining shovel as the front end loaders become appropriately larger and more reliable. There has been some trend in this direction in the past with smaller shovels and loaders. In Table 1 are data for U.S. production of front end loaders (5 to 6-cu-yd class and 10-cu-yd class) and U.S. production of cable-operated, crawler-mounted power shovels, draglines, and cranes (11/z to 5-cu-yd class) for the years 1965 and 1970 (Source: Bureau of Census, U.S. Department of Commerce). Many believe that the decline in shovel production, in particular, is due at least in part to the acceptance of the front end loader as a primary production tool. However, the foregoing comparison is for a smaller shovel class than for front end loaders. It has generally been a rule of thumb that a front end loader is competitive on a cost and production basis with a shovel when it is approximately twice as large as the shovel. This same relationship may not hold true with the larger 10 to 15-cu-yd electric shovels, but at some point, direct competition will probably occur. Current Trends There are many trends today that are significant. One major usage trend in large front end loaders involves load and carry operations. This application is defined as "performing the functions of loading, hauling and dumping with one machine over distances within its economical limits." Many people are using front end loaders to transport material as well as load it. It is estimated that nearly half of all 10-cu-yd front end loaders sold in 1970 went into primary load and carry work. Many times, significant savings result from using the load and carry system, instead of small truck and shovel operations, for distances in excess of 1000 ft. Strip mining in coal utilizes great numbers of front end loaders in this manner. In-pit crushing of limestone is developing on an ever-increasing level, especially in Europe, and this stimulates interest in load and carry operations. In Table 2 is a cost example for 10-cu-yd front end loader in a limestone quarry. More than likely, load and carry use will continue to increase over the next few years. While some types of mining do not lend themselves to load and carry in the pit, such areas as surge piles are applicable and should be studied. In the area of machine operating cost reduction, there

Jan 1, 1974

By W. T. Lundy

THE forms in which sulphur is commonly found-native sulphur, sulphides of many metals and sulphates-are widely distributed throughout the world. The two first mentioned are the principal sources of sulphur. Substantial amounts of sulphur in some form are recovered from gases from industrial plants and from gases from smelters treating other minerals. In recent years, plants have been installed in the United States to recover sulphur in elemental form from sour gases."' The amount of sulphur so produced is small at the present time but undoubtedly will increase in the future. The recovery of sulphur from gases is not covered in this chapter, which is confined to: (1) native sulphur deposits of Texas and Louisiana; (2) those of Italy; (3) other scattered deposits of native sulphur; and (4) pyrites deposits found in many countries of the world.30 Sulphur, a nonmetallic element, occurs normally in well developed bipyramidal or tabular crystals of the orthorhombic system, as well as in stalactitic and earthy masses. The mineral has a distinctive yellow color, which may be dark and discolored by impurities; the streak is white; the luster is resinous; and the fracture is conchoidal to uneven. Sulphur is translucent to opaque; has a hardness of 1.5 to 2.5, a specific gravity of 2.05, and an atomic weight of 32.06.16 The molecular weight of the solid is expressed by the symbol SS and that of the liquid and vapor ranges from S2 to S8, depending on the temperature. Sulphur melts at a temperature between 2340 and 248°F, depending on its crystalline state, to a liquid of dark amber color, the boiling point of which is 832°F. The viscosity near the melting point is about that of water and increases to 50,000 times that of water at a temperature of 380°F, subsequently falling rapidly to a low figure at the boiling point. Crystallization changes slowly from rhombic to monoclinic above 205 °F and in cooling the transformation is reversed. The mineral is a very poor conductor of heat and electricity; it is insoluble in water and nearly all acids and is soluble in cold carbon bisulphide and carbon tetrachloride. It ignites in air at a temperature of 478°F, burning with a distinctive blue flame with evolution of sulphur dioxide.47

Jan 1, 1949

By Howard Eavenson

IN a discussion in the Transactions of the Institute (vol. xl, page 835 et seq.) the writer gave some data about explosions of gas and dust in the coal mines of the United States, Canada, and Mexico, among which was a table showing all explosions in the countries mentioned, in which five or more fatalities occurred, which had been reported from Oct. 2, 1871, to Mar. 28, 1908. Since the elate of publication of this discussion, as opportunity has presented, this table has been extended by the addition of explosions occurring since that time, by including several explosions which occurred prior to 1871, and by adding to the list all explosions of gas or dust in which less than five fatalities occurred, this having been clone by a search through the various mine reports, partly by the author and partly by the Librarian of the United Engineering Societies, and by correspondence with mine inspection officials of the various States, to all of whom acknowledgment of assistance is made: The accompanying list, Table I, is believed to be as nearly correct as such a statement can be made, and, while only extending to Dec. 20, 1912, contains all of the data now available in reports. An analysis of Table I, showing the occurrence by months of all explosions of gas or dust, including those in which less as well as those in which more than five fatalities occurred, is given in Table II. The statements are frequently made, in technical papers, that the number of explosions, particularly the larger ones, is increasing more rapidly than the number of mines, and that the number of fatalities caused by mine explosions, particularly the more disastrous ones, is increasing faster, proportionally, than the production of coal. To investigate the correctness of these statements Table III was prepared, in which the number of accidents and fatalities is taken from Table I, using those in the United States only, on account of lack of other data for Canada and Mexico; the production is from the records of the U.S. Geological Survey, and the number of mines, from 1909, from the same source; previous to 1909, the number of mines, for the decennial periods, was furnished by the U. S. Census records.

Jan 10, 1914