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By Richard R. Wilding

The United States Coal Mining Indus- try had the opportunity to greatly improve their roof control techniques during the nineteen seventies by the use of fully grouted resin roof bolts. This new tool was first approved for underground coal mines by M.S.H.A. ten years ago and during the next decade has been perfected to meet many problems previously thought to be unsolveable. The initial approved cartridge was very similar to the ones used in Europe in the nineteen sixties and since then has undergone a great deal of modification for our mines. These refinements in cartridge construction have been obtained by the close cooperation of resin manufacturers with mining personnel in perfecting pro- ducts customized to particular needs. This type of continuing program naturally necessitates a responsibility upon the resin anchor manufacturer to maintain a competant technical department. These roof control technicians by constant exposure to many of the most demanding underground problems, have developed through the years into leaders in successful applications. Today, Celtite has over twenty technically trained underground technicians with an accumulation of over one hundred years experience in resin anchors. Our documented technical reports are in excess of five thousand, involving over six hundred mines representing every major coal mining area in most mineable seams. This department has been the implement used to work with our research and development people in perfecting a continual line of improved resin anchors. This program has been a significant factor in leading the resin anchor industry into a position where approximately one third of the 1980 under- ground coal production used the fully grouted system. However, as gratifying as this accomplishment is, our experience has lead us to believe that there is an additional demand for a roof control system capable of combining the mechanical bolts ability to apply tension with resins proven superiority in anchoring. A significant number mines visited during the last decade had problems that we believed could be better answered by applying tension rather than the fully grouted approach. In most in- stances, their current mechanical installations were doing an adequate job until anchor slippage occured. Fully grouted systems solved this bleed-off problem. However, an acceptable number of falls still occured. In these some- what unique areas which we believe may amount to as much as fifteen percent of the total bolt installations, the need for improved anchors was a necessity. In developing this program we listed the ultimate requirements necessary for both the bolt and the anchor. These are for the bolt: A. Have one end manufactured so it can be grouted to the mine roof with polyester resin, obtaining a variable length from 12" to 36". B. Have a mechanism which allows the operator to tension the roof from zero to 10,000 psi. This mechanism should indicate if the roof is in compression and the bolt in tension. C. Have a mechanism which allows the operator to thoroughly mix the resin. D. The head of the bolt should not extend below the roof line any more than a normal mechanical bolt. E. Not require any extra tools for installation. F. Be bendable for use in low seam coal. G. Be durable and inexpensive. The ultimate point anchor resin should have the following characteristics: A. Have a mix and gel cycle which would compete with a mechanical bolt for speed of installation. B. Have a pull out strength in excess of a normal 5/8" mechanical bolt.

Jan 1, 1981

By C Quirion

Agnico-Eagle Mines Limited, Laronde Division, operates the deepest single-lift shaft in the western hemisphere. The initial Penna hoisting plant was designed for a production rate of 1800 tonnes per day (t/d) from a 2240 m deep shaft. Exploration success, while sinking the shaft, led to a decision in 1998 to increase the production rate to 4500 t/d of ore from a loading depth of 2200 m. The new production rate was achieved by increasing the size of the rope and reducing the rope static factor from the existing 5.0 required by regulation to a static factor using the formula 25 000/(4000 + L) contained in a draft document that would eventually become SABS 0294 in South Africa. The formula permits Laronde to use a static factor as low as 4.0. The upgraded Penna hoisting system is the first hoisting plant outside of South Africa to meet the South African requirements. This paper describes the conversion of the hoisting plant to the lower rope selection factor in the year 2000 and the operating and maintenance issues experienced since then.

Jan 1, 2005

By Nicholas J. Grant, Klaus M. Zwilsky

EVER since the unusual high temperature creep resistance and structure stability of SAP (Sintered Aluminum Powder) and similar aluminum-alumina alloys were reported,'," there has been a need to determine whether other metal-metal oxide (or other hard stable particle) alloys, not so ideal as the Al-Al 2 O 3 system, would show similar high temperature strength and stability. Of the many possible methods of obtaining a metal-metal oxide alloy,a one that offers many practical and theoretical advantages is the simple mechanical blending of metal powders with hard stable particles, followed by pressing, sintering, and hot extruding. With this purpose in mind copper powder and silica and alumina were selected for experimentation because these materials are readily available, cheap, reliable, and lend themseives te the simplest processing and testing. Starting Materials—Two sizes of electrolytically deposited copper powders were used, a —200 mesh powder referred to as A powder in this investigation and a —325 mesh powder referred to as C powder. Silica was added in the form of pure Potter's flint from which a —10µ fraction had been separated. Alumina of the a variety was used in two size ranges. The first was a —10µ fraction separated from a coarser aggregate, the second a commercial product of 0.3µ average diam manufactured by Linde Air Products. Preparation of Powders—Five-hundred gram batches of A copper with 1, 3, and 10 volume pct of each of the three oxides mentioned above were prepared, as well as C copper containing 1, 3, and 10 volume pct of —lop silica. Powders were weighed out carefully and blended by ball milling in air for 24 hr. Since this treatment resulted in surface oxidation of the copper, the charge was hydrogen reduced at 260°C for 5 hr. The next step consisted of hydrostatically compacting the powder into a shape suitable for the required hot working. For this purpose a slug about 11/2 in. in diam and 31/2 in. in length was produced by placing the mixed powders into a rubber sleeve, holding in a perforated steel canister, and rubber stoppering at the ends; the assembly was then subjected to a hydrostatic pressure of 50,000 psi. The air was evacuated from the assembly prior to pressing in order to prevent

Jan 1, 1958

Jan 1, 1931

By Paul M. Tyler, R. M. Santmyers

Platinum and its allied metals osmium, iridium, rhodium , palladium, and ruthenium are the most costly of the better-known metals and combine properties that make them unique among the elements . For a number of isolated uses substitutes have been found which are cheaper and quite as satisfactory ; but none will replace platinum in many of its varied uses , and there is a large and growing demand for purposes for which no economical substitutes are known . As this paper is being written the supply of platinum tends further to increase as a result of the exploitation of newly developed resources and the expansion of production from older sources . In obedience to economic law prices have declined steadily . Buyers , always reluctant to commit themselves on a falling market , have maintained a waiting attitude , and consumption trends are not clearly indicated . Should platinum again approach the price of gold , the natural tendency would be to use more of the metal . In recent years , however , the industrial and scientific uses of platinum have become subordinate to its consumption in jewelry . In matters of personal adornment the popular whim and not practical utility is the dominant factor in demand , and it does not follow that lowered prices will stimulate demand . The substitution of white gold and certain plated goods for platinum may be checked by cheaper platinum , but a fickle public that turned from gold to platinum when platinum became the more costly metal can not with certainty be expected to buy more platinum jewelry because more of the metal than formerly may now be had for the same price . The dearth of platinum during and for some years after the World War had a profound effect upon prices , and this in turn wrought important changes in the field of consumption . When prices advanced the chemical industry, formerly the leading consumer of platinum, was forced to economize in its consumption of the metal and to substitute other metals whenever it could . Somewhat similar conditions resulted in the electrical industry and in the use of platinum in dentistry . Some of these former outlets are permanently closed , and others may be reopened now that platinum prices have receded to pre-war levels , but despite the general expansion of industrial and scientific activity since 1913 the demand for platinum outside the jewelry trade has suffered a net decline . Since the metallurgy of platinum is relatively simple the reason for its high price is its relative scarcity in the earth's crust . An interesting speculation set forth by certain geologists attributes the extraordinarily high density of the interior of the earth in part to the presence of iron alloyed with substantial amounts of platinum and other related metals . In support of this hypothesis it may be said that platinum is often found in meteorites , and known deposits in the earth's crust are all of deep - seated origin . However , even if there were a high concentration of platinum in the interior of the earth it would be next to impossible at the present state of our knowledge to make it available , and the world would continue to depend for its supply on deposits found within minable depths below the surface . For many years before 1914 the world obtained its supply of platinum metals primarily from Russia , which before the war produced as much as 300,000 troy ounces in a year , or approximately 95 per cent of the world output . Despite fluctuations due to internal political disturbances Colombia managed to account for 3 to 4 per cent , leaving the rest of the world to furnish the remaining 1 to 2 per cent .

Feb 1, 1931

By G. L. F. Powell, L. M. Hogan

Large degrees of undercooling have been produced in bulk samples, 400 g, of copper and Cu-O alloys by melting in a slag of commercial soda-lime glass. The maximum degrees of undercooling obtained for copper, hypoeutectic and hypereutectic Cu-O alloy samples were 208°, 218°, and 97°C, respectively. No grain refinement as a function of undercooling was observed for the pure copper samples although in Cu-O alloys there was a marked decrease in grain size at degrees of undercooling greater than 150°C. The grain size change is the result of recrystallization during or immediately following the freezing process. THE first report of a large degree of undercooling in a bulk sample of metal was that by Bardenheuer and Bleckmann1 who produced 258°C undercooling in a 150-g sample of iron by melting in a glass slag. Garbeck2 extended this technique to nickel, cobalt, and copper and reported maximum undercoolings of 220°, 230°, and 60°C, respectively, while Fehling and schei13 undercooled a large number of metals as 2 to 20-g samples slagged in glass. Subsequently, one of the authors4 applied the glass slag technique to bulk samples, -500 g, of silver and obtained a maximum undercooling of 250°C. It was concluded that impurities which normally catalyze nucleation at small degrees of undercooling could be removed by oxidation and solution in a glass slag. Large undercooling of bulk samples of iron, nickel, cobalt, and their alloys has also been reported by other authors.8'10'11 Of the five metals mentioned above, copper was the only one which had not been undercooled by at least 200°C in the form of a large bulk sample. Since copper exhibits high liquid solubility for oxygen, as do iron, nickel, cobalt, and silver, it was considered that a much larger degree of undercooling than that observed by Garbeck for copper should be possible. This was obtained by modification of the technique applied previously to silver and the results of undercooling experiments with 400-g samples of copper form part of this paper. The undercooling behavior of Cu-O alloys was also studied, and the influence of a small oxygen content on the grain structure of undercooled copper was observed metallographically. 1) EXPERIMENTAL The copper was obtained as oxygen-free high-conductivity copper, the major impurities of which were Fe 0.01 pct, Zn 0.015 pct, and Si 0.02 pct. Melting was carried out in an open-ended vertical cylindrical fur- nace wound with Kanthal wire. A schematic diagram of the experimental setup is shown in Fig. 1. The undercooling technique previously used for silver4 involved preoxidation of the samples by melting the silver in contact with the atmosphere. The oxygen content was subsequently reduced by freezing the samples under a glass slag, whereby the oxygen was evolved as gas and the continuous glass cover over the surface prevented re-solution of oxygen when the sample was remelted. Since oxygen is not released as gas when a Cu-O alloy is frozen. the technique had to be modified for use with copper. Also, copper is oxidized during fire refining so that an in situ preoxidation treatment of the samples was not considered to be a prerequisite to large undercooling. Experimentation proved that this supposition was correct. Instead, it was found that care was necessary during melting of the copper under glass to minimize oxygen pickup. Samples of copper weighing 400 g were prepared by adding 50-g pieces of copper to a vitreous silica crucible partly filled with soda lime glass at a temperature of approximately 1000°C. Each piece was quickly immersed in the glass slag and held at this temperature until the oxide coating on the surface decomposed. This change was easily observed and coincided with the formation of gas bubbles in the glass

Jan 1, 1969

PRESIDENT MOSES: I thank you, Mr. Emrick. Mr. Shubart would like to supplement your paper with another article. SECRETARY SHUBART: My paper really should not supplement Mr. Emrick's paper, because I secured some data from him, I knew Mr. Emrick had had a great deal of experience in shearing and I thought he could give you some useful data. (Here Mr. Shubart read the following:) "For the past two years, there has been considerable discussion among coal mining men concerning the advantages of a shearing cut in addition to the undercut, in mine work. We have had a number of requests for information on sheared coal. We were asked questions like these: What percentage increase in lump can be obtained from sheared coal? What has been the experience of mine men who have employed the shearing cut? Is the shearing cut a means of increasing the lump content? If so, is it increased to such an extent as to warrant the installation of shearing machines? Upon investigation it was found that no one could give us any definite information on the subject as no one had weighed separately the grades of coal produced from' the shearing cut. With a view of getting this information first-hand, the Goodman Manufacturing Company built up a shearing machine to be used in experimental work only. The officials of the Troy Coal Company were eager to share in the expense of obtaining this information, so that in May of this year we shipped to their mines our first type 319 shearing machine. The Troy Coal Company operates a mine at Troy, Illinois, in the No. 6 seam and produces about 275 tons daily. The coal seam averages five and one-half feet in thickness. Previous to the arrival of the shearing machine the coal was undercut six feet deep by a' Goodman Shortwall Machine. To break down the coal required two shots in the entries and three in the rooms. The percentage of lump measured over two-inch screens, averaged 58.2 per cent. The conditions of the mine were very favorable for obtaining accurate weights on the amount of lump and slack coal run over the shaker screens. The shearing machine easily cut the entire tonnage of the mine so that all coal dumped into the railroad cars had been produced from a shearing cut, in addition to an undercut. Accurate records were kept of the railroad weights and front this data the percentages of lump and slack were computed. During the month of June the percentage of lump produced over two-inch shaker screens was 69.9 per cent. This result demonstrated that the introduction of the shearing cut had increased the percentage of lump twenty per cent. The entries were driven ten feet wide and the rooms thirty feet wide The shearing cut was made near the center of the face in all places. A hole was drilled in each rib and the coal shot down with one-half the powder required when the places had been undercut only. The coal was broken down to the back of the cut so that very little pick work was necessary to load it out. The benefits derived from the shearing cut were as follows: 1. Increased percentage of lump, thus increasing the average price of coal per ton. 2. Preservation of a tender top due to the decreased amount of explosive used. 3. Facilitated handling of coal by the miner thus increasing the tonnage per man. 4. Eliminated coal yardage in entries and crosscuts of ten feet or less in width for the reason that the shearing cut provides a loose end from which to shoot the coal. This applies to the union fields only, as the non-union mines do not pay for coal yardage. Based on a price of $2.50 for lump coal and $1.25 for slack coal, the average price per ton was increased

Jan 1, 1926

The MineSight Schedule Optimizer (MSSO) is a new production scheduling tool developed by Mintec, Inc. using mixed integer linear programming (MILP) techniques. This tool was developed to solve the short- and medium-term schedule problems for surface and underground mining operations that involve multiple mining areas, multiple processes, multiple destinations and blending requirements. MSSO finds the optimum schedule in each period that achieves the objective while satisfying comprehensive product quality and quantity requirements as well as physical and technical constraints. Using the MineSight Planning Database (MSPD) as a central data hub, MSSO seamlessly works with MineSight 3-D (MS3D) block models and third-party data, MineSight Interactive Planner (MSIP), MineSight Haulage (MSHaulage) and MineSight Activity Planner (MSAP) to prepare data, formulate MILP models, calculate, report, fine-tune, save and visualize schedules. This paper presents the features and capabilities of MSSO. A case study is also given to demonstrate the applications of this new tool.

Jan 1, 2009

By D. J. Bauer

Oxidation of mixed lanthanide solutions with ozone at pH 4.5 and at ambient temperature resulted in precipitation of 98 percent of the original cerium present. Corresponding cerium purity was increased from 50 to 98 percent. A second ozone oxidation-precipitation step produced cerium of 99.9 percent purity in high yield from the 98-percent-pure material. Alternately, ozonation at pH 6.5 and 85° C differentially precipitated cerium and rare-earth elements heavier than cerium and left 89 percent of the original lanthanum in solution at a purity of 95 percent. Subsequent recovery of cerium from the heavier rare-earth elements in the precipitate was accomplished by dissolving the precipitate in dilute mineral acid and reoxidizing the cerium with ozone at pH 4.5 and at ambient temperature. The filtrate contained an enriched praseodymium-neodymium-samarium-europium mixture that is amenable to separation by ion exchange or solvent extraction.

Jan 1, 1968

By M. J. Fairweather, Keith Lee, Naseem Mian

"The initial metallurgical response of the lead flotation circuit at Cominco Alaska's Red Dog Mine was substantially below feasibility expectations. A program of diagnostic mineralogical analysis pointed the way to a number of process improvements, including simplification of the flow sheet and installation of a large flotation column as a lead cleaner. A diagnostic mineralogical analysis of the improved circuit shows that inadvertent activation of both sphalerite and pyrite has been substantially reduced, resulting in significant improvements in metallurgical performance. Directions for further improvements to the circuit performance are suggested by the recent analysis.IntroductionRed Dog is a high-grade, sedimentary exhalative (sedex) deposit with relatively simple mineralogy. The primary non-sulphide gangue minerals are quartz and barite, with minor barium carbonate (witherite). The iron sulphate minerals, siderotil, rosenite and copiapite occur variably throughout the deposit. Mineral texture is variable ranging from liberated sphalerite grains visible to the naked eye to intergrowths, particularly between sphalerite and galena, to less than ten microns. Both pyrite and marcasite are present. To date no enhanced reactivity of marcasite has been observed. For this discussion, ""pyrite"" will be used to refer to both iron sulphide minerals. In common with other sedex deposits, Red Dog contains organic and sulphur impurities."

Jan 1, 1997

By D. P. Richards

Tunneling in carbonate sedimentary rocks such as lime stones and dolomites can be difficult when solution channels or voids are anticipated. Addressing the risk aspects associated with these formations can be done effectively through the Geotechnical Baseline Report (GBR) by properly characterizing the ground and ensuring that the Owner?s and the Contractor?s interests are represented and protected. This paper presents a review of tunneling difficulties in these rocks and includes a review of site investigations, geologic characterization, and presentation in the GBR. Key construction issues are also discussed, including appropriate excavation technology, mechanisms of face support and groundwater control, and quantifying grout takes.

By N. Shariatmadari

Two dry docks are currently under construction on the northern banks of the Persian Gulf by Iran Shipyard and Offshore Industries Co. (ISOICO). The construction of these dry docks forms a huge civil offshore project. The foundation system include more than 6500 large diameter (40-48 inch) spiral pipe piles with embedded depth varying from 18 to 25 meters. Extensive geotechnical subsurface investigations were carried out in the project site in order to provide enough information and proper perspective on the underlying soil layers. The results of geotechnical investigation indicated that a 20 m thick dense silty sand layer exist which can be considered as the main bearing stratum. Due to short clear distance between the piles (less than 3 m) refusal occurred for some piles, accordingly led to actual length less than their design length. To verify and in-situ evaluation of piles capacity, static and dynamic full scale loading tests programme were performed. The data from 40 static compression and tension and also 50 dynamic tests with PDA were interpreted and analyzed by current approaches and CAPWAP software. The results of analysis demonstrated that the safety factor of bearing capacity of piles must be increased to reach a safety factor of about 1.8. To achieve this purpose the interaction of combined piled-raft system has been realized.

Jan 1, 2007

By M. Ounpuu

"Gold, because of low head grade and recovery, is usually of minor importance in the economics of base metal flotation mills. Because of the above little metallurgical attention is usually provided to gold in these mills. But collectively, the gold losses can be significant. For instance, Noranda presently has interest in 13 base metal flotation mills with a collective gold recovery of only 50%. Gold losses are in the order of 150.000 oz per year!The Noranda Technology Centre, in cooperation with Noranda's milling operations, has been investigating the potential of gravity concentration of gold from base metal flotation mills. This paper outlines the pilot and plant testing conducted at three operations with distinctly different ore types; porphyry copper, complex massive sulphide and polymetallic semi-massive sulphide. Testing has been with the Knelson Concentrator which has proven to successfully recover fine gold in a simple, trouble free circuit. Discussion of issues which have risen from the testwork is included. INTRODUCTIONGold, because of low head grade and recovery, is usually of minor importance in the economics of poly-metallic base metal ores. TABLE I presents gold heads and recovery for some of Noranda's base metal mills. It should be noted that the stated recoveries are to payable concentrates, but does not reflect the actual payable gold recovery which in cases is significantly lower. Noranda's average gold recovery into a gold payable concentrates is about 45%. This equates to a collective loss in the order of $57 MM per year of gold"

Jan 1, 1992

By Greg Rasmussen

"Huckleberry Mines Ltd is an open pit porphyry copper-molybdenum mining operation located 120 km south of Houston, British Columbia with a 18,000 tpd milling facility. Huckleberry Mines is owned 60% by Princeton Mining Corporation, the operator, and 40% by Japanese partners.The Huckleberry deposit is one of several porphyry systems located in the Tahts a district of the intermontane belt of British Columbia. The principal economic minerals are chalcopyrite and molybdenite. The ore deposit is split into two ore deposits, the Main Zone, consisting of24,241,000 tonnes of mineable reserves grading 0.484 %Cu and 0.013 %Mo, and the East Zone, consisting of 66, 131,500 tonnes of mineable reserves grading 0.523 %Cu and 0.014 %Mo.The milling begins with gyratory crushing. This sizes the product down to 100% passing 6 inches. This product is discharged onto a stockpile by a stationary conveyor. The stockpile was designed for a capacity of 144,000 tonnes with a live capacity of 20,000 tonnes. Two apron feeders draw the ore from the bottom of the stockpile and discharge onto a SAG mill feed conveyor which in turn feeds a grinding circuit consisting of one 32' x 15' SAG mill and two 30' x 16.5' ball mills. The grinding cyclone overflow feeds bulk flotation at 55 -65% minus 200 mesh.The bulk flotation uses conventional rougher and column flotation methods to float a bulk concentrate recovering 92.6% copper on the East Zone and 95% copper on the Main Zone with 75% recovery molybdenum on both zones. The bulk concentrate feeds the bulk thickener which produces a density of 60 %solids. This in turn feeds the molybdenum circuit, which separates the copper and molybdenum, by depressing the copper then using conventional rougher and column flotation produces a copper concentrate of 27 %Cu and 0.02 %Mo, and a molybdenum concentrate containing 1.5 %Cu and 54 %Mo at 67% recovery. These concentrates are separately thickened, filtered, dried, and shipped to the smelter for processing."

Jan 1, 1998

By James N. Bayne

An examination and analysis of Landsat digital data in the monitoring of surface mines in selected counties of South Carolina.

Jan 1, 1979

By Terry Keough

Contents for discussion ?Current liner technology ?Innovation of magnetic products ?Development of magnetic wear liners ?Installation and change out method ?Uses and capabilities ?Internal wear liners ?Summary Vulcanized rubber ?Many of you are familiar with using vulcanized rubber to provide wear protection on mining equipment. For years, this was the standard for most mining equipment, and mines spent a lot of money fabricating and lining steel equipment.

May 1, 2010

By Albert B. Horn, J. J. McKetta, O. L. Culberson

Since water is present in natural gas and petroleum reservoirs, it is of engineering value to have accurate experimental data regarding the behavior of water in hydro-carbon systems. Since experimental data of this nature are so scarce in literature, a program is under way at the University of Texas to obtain experimental data regarding the distribution of water in petroleum hydrocarbons. This initial paper is subdivided into three parts: A. The apparatus and analytical procedure are discussed. B. The equipment has been checked by comparing the experimental data with data of literature on the solubility of methane in water at 77°F and pressures to 10,000 psia. C. Experimental data are presented at 100, 160, 220, 280, and 340°F on the solubility of ethane in water at pressures to 1200 psia. As the temperature is increased the solubility of ethane in water first exhibits a minimum and then a maximum solubility. The range of applicability of Henry's Law is discussed. INTRODUCTION Although considerable experimental data have been published on water-hydrocarbon systems in the hydrate region, the information on these systems in the vapor-liquid and the vapor-liquid- 'References are given at end of paper. Manuscript received at the office of the Petroleum Branch August 31. 1949. liquid regions at elevated pressures is extremely scarce. A survey of the literature on hydrocarbon-water systems at elevated pressures was presented by McKetta and Katz12,13,14 showing the systems and phases investigated. In addition to these existing published data, Poettman and Dean" have reported the water content of propane in the three-phase region (only the propane-rich liquid and the vapor phases were studied) and Michels, et al.15 have reported the solubility of methane in water at pressures to 6800 pounds per square inch at 77°F and at pressures to 1500 psia at temperatures to 300°F. It has been shown12,14 that the presence of water will materially affect the phase equilibrium of hydrocarbon systems. In order to make experimental studies on the quantitative effect of the presence of water in naturally-occurring hydrocarbon systems, it is desirable to have data available on binary hydrocarbon-water systems. A program is underway at this laboratory to determine the phase relations of the hydrocarbon-water systems. Initially the investigation will be limited to binary systems with water as one of the components. Later ternary and complex systems will be investigated and reported. EQUIPMENT AND ANALYTICAL PROCEDURE Apparatus In order to avoid a description of the methods employed in each subsequent paper, a description of the apparatus and the analytical procedure will be given here. The apparatus is shown schematically in Fig. 1. The numbered items in Fig. 1 represent the high pressure valves. The equilibrium cell is of the totally enclosed type, with a volume of 550 cubic centimeters. The cell is mounted in an air bath which in turn is mounted on pillow blocks and connected to a rocking mechanism so that the cell and bath will oscillate approximately 30 degrees above and below the horizontal at 35 oscillations per minute. The air bath is heated by two resistance heaters. The first or primary heater is 250 watts and is controlled by a thermoregulator; the second heater is of 500-watt size and is manually controlled to aid in bringing the cell to equilibrium temperature. Air circulation in the bath is accomplished by means of a fan. the shaft of which is concentric to the heaters. The walls of the bath are insulated with 2-in. thick blocks of 85 per cent magnesia. Temperature observations are made with three calibrated thermocouples. One couple is sealed in a well provided within the equilibrium cell, and is used as the criterion of the cell being at the equilibrium temperature. The other two thermocouples are located on opposite sides of the cell to indicate temperature uniformity within the air bath. The thermoregulator, used in conjunction with a relay, gives a variation in the temperature of the air in the bath of plus or minus one degree Fahrenheit. However, since the equilibrium cell weighs approximately 55 pounds, the

Jan 1, 1950

By V. Koump, T. F. Perzak, R. G. Olsson

The rate of dissolution of recrystallized alumina in molten iron oxide in equilibrium with iron at 1450°C was studied by rotating alumina disks in the melt. The samples were rotated from 1 to 12 min at speeds from 54 to 3270 rpm. It is concluded from the experimental results that the process is limited by diffusion in the liquid boundary layer. The interdiffusion coefficient for the dissolution of alumina disks in pure molten iron oxide was estimated to be about 3 xl0-5= sq cm per sec. THE erosion of refractory materials is of considerable importance in iron- and steelmaking processes. Since many refractories contain alumina, this investigation was undertaken to develop a further understanding of the erosion mechanism. Shurygin et al.1-3 have investigated the rate of dissolution of alumina in molten silicates and molten fluorides by rotating alumina disks in various melts. The rates were reported to be limited by diffusion in the liquid boundary layer. In the present investigation the rate of dissolution of alumina in molten iron oxide in equilibrium with iron was studied in a similar manner; alumina disks were rotated at different speeds in molten iron oxide at 1450°C. The relationship between the rate of dissolution of the flat surface of the disk and the speed of rotation was used to determine whether the process is limited by diffusion in the liquid boundary layer or by some chemical reaction mechanism. APPARATUS AND EXPERIMENTAL PROCEDURE A schematic diagram of the experimental apparatus is shown in Fig. 1. The iron oxide melts were contained in 2-in.-diam high-purity iron crucibles that were centered in a cylindrical graphite susceptor. Prior to each sequence of experiments, sufficient reagent-grade ferric oxide was added to a crucible to form approximately 350 g of molten iron oxide; the depth of the melt was about 11/2 in. To prevent excessive erosion of the crucible, about 70 pct of the required iron was added to the crucible in the form of iron powder. The mixture was melted in an atmosphere of purified argon by induction heating with a 250-kc generator. After the oxide was molten, a Pt/Pt-10 pct Rh thermocouple at the exterior of the crucible was calibrated against an iron-sheathed thermocouple immersed in the melt and was thereafter used to regulate the melt temperature. All experiments were conducted at 1450°C with a probable error of 10°C. The alumina disks were constructed from pure re-crystallized alumina having a density of 3.7 g per cu cm (porosity approximately 6 pct). For each disk sample a 1-cm-diam alumina cylinder of known length (0.9 to 1.0 cm) was cemented into a ground seat at the end of a 1.27-cm-OD centerless ground alumina tube. The ends of the tube and cylinder were flush and served as the disk surface. When mounted in the apparatus the alumina tube was guided by tungsten carbide bearings and rotated by a variable-speed motor. The bearing and drive assembly could be raised and lowered in order to allow the sample to be quickly moved in and out of the melt. At the start of an experiment, a sample was preheated for several minutes directly above the melt, and then immersed in the middle of the melt to a depth of a in. and rotated for 1 to 12 min at speeds of 64 to 3270 rpm. The speed of rotation was measured by a hand tachometer. At the start and completion of each experiment, a melt sample was withdrawn with a cold copper rod and the alumina content was determined by chemical analysis. During individual experiments the alumina concentration of the melt changed by the order of 0.25 wt pct. Average melt compositions for individual experiments were used in the subsequent computations and were in the range between 0.12 and 4.3 wt pct alumina. The extent of dissolution was determined by sectioning the end of the sample and measuring the final length of the alumina cylinder with a microscope. Since the exposed end of the cylinder, i.e., the disk surface, was etched in a manner characteristic of this geometry, a number of measurements were made on

Jan 1, 1969

By David S. Bolin

INTRODUCTION This paper is directed toward those companies or individuals who may be considering the possibility of tin exploration or development projects in South America. Although tin deposits are known in many countries of Latin-America including Argentina, Peru, and Mexico, the majority of the deposits are located in Bolivia and Brazil. These two countries also account for virtually all the current production. Many factors affect the economic decisions related to mining and exploration projects in this region including the following: 1) Types of deposits 2) Anticipated size and grade of deposits 3) Deposit geometry and ore distribution as it affects the selection of a mining method 4) Metallurgical amenability 5) Governmental policies 6) Taxation 7) Anticipated capital and operating costs 8) Marketing costs This discussion will be directed toward each of these points. The majority of the presentation will be concentrated on Bolivia as this country is the principal producer in the region, however, the potential for further tin development in Brazil is excellent. Due to the remote and previously almost inaccessible location of the stanniferous districts of Brazil, little is known with respect to size and type of non-alluvial deposits which may exist in this vast country. TYPES OF DEPOSITS Two major types of deposits are currently being exploited in Bolivia; alluvial, and hard rock or lode deposits. Bolivia produces substantial tin from both types of deposit whereas virtually all Brazilian production to date has been from alluvial sources. Alluvial Deposits Brazil: The alluvial tin deposits of Brazil are located in river channels and flood plains adjacent to low mountain ranges. The terrain containing the tin placers is flat, marshy, and generally jungle covered. The major controls of alluvial cassiterite concentration are the ancient and present stream channels. The average tin concentration in the placers varies from 500 grams to approximately 1.0 kilograms per cubic meter. Tin reserves in the Rondonia field of Brazil have been estimated at 600,000 tons of fine tin. A bucketwheel suction dredge went into production in the Rondonia district in 1979, and four others have since been ordered. Several other gravel pump, and hand mining operations are also in production in this field. In addition to the Rondonia district, tin occurrences are known from Xingu, in Para state, and in the state of Minas Gerais. Bolivia: The alluvial deposits of Bolivia are somewhat more complex due to the variable geomorphology and abrupt topography. Conventional placer accumulations of cassiterite are found in many stream channels and intermontane basins surrounding the major lode tin producing regions. In addition to stream and valley placers, a group of deposits locally referred to as "Pallacos" or "Llamperas" which consist of colluvium, landslide debris and glacial moraine material, contain substantial tin reserves in some areas. The stream channel and intermontane basins contain the only deposits which are presently being exploited by mechanized methods. One dredge is working the stream channel below Cerro Rico de Potosi and another is operating in an intermontane basin southeast of the city of Oruro. Both of these dredges are operated by private companies. The average grade for these operations varies from 250 to 500 grams per cubic meter. The largest of the intermontane basin placers known at present is the Centenario deposit located adjacent to the Catavi lode deposit. This deposit contains approximately 170 million cubic meters of material with an average grade of about 150 grams per cubic meter. The "Pallacos" deposits are found on the slopes of mineralized areas and in glacial moraine. The mineralized material is generally completely unsorted, with tin and sometimes tungsten values distributed erratically throughout the entire mass. Most of these deposits are worked by small leasors or cooperatives; however, at least one mechanized washing plant is in operation southeast of Oruro. The size of these deposits may reach up to several million cubic yards. Grades are very erratic, but may range from 200 to 500 grams per cubic yard. In addition to the formal mining operations, virtually every drainage surrounding the major mines is being worked by independent' miners utilizing hand mining and jig or pan concentration. The aggregate production from these operations is substantial. The

Jan 1, 1982

By Barton J. Suchomel

Introduction Despite perennial economic turbulence, Argentina continues to push ahead with efforts to develop its fledgling hardrock mining industry. The country lacks the infrastructure and mining history of neighboring Chile, but contains vast areas of great resource potential. The eastern slope of the Andean Cordillera in west-northwest Argentina is a relatively little-explored region offering geology highly conducive to the occurrence of base and precious metals deposits. Past exploration has concentrated on locating and evaluating porphyry copper-molybdenum (Cu-Mo) type occurrences. Comparatively little attention has been given to precious metals potential. An interesting and important group of base and precious metals prospects in northwestern Argentina occurs within and adjacent to the Upper Miocene Farallon Negro volcanic complex in Catamarca province. The volcanic complex contains numerous base and precious metals vein occurrences, high gold porphyry copper prospects, and possible disseminated gold targets that have attracted relatively minor attention in the past. Various geologic aspects of the region have been reported by several workers (Llambias, 1970, 1972; Caelles et al, 1971; and Sillitoe, 1973). Regional geology The Farallon Negro volcanic complex is located within the Sierras Pampeanas morphostructural province about 200 km (125 miles) east of the main Andean Cordillera (Fig. 1). The Sierras Pampeanas is characterized by north-south-trending basin and range topography that developed during the Miocene and Pliocene. Uplifted blocks are composed mainly of lower Paleozoic metasediments and granitoid plutonic rocks unconformably overlain by Tertiary redbeds and minor carbonates. A consistent landward migration of volcanic and intrusive ac¬tivity occurred in this region of the Andes from the early Mesozoic through the mid-Tertiary. An eastward extension of magmatism during the Miocene and Pliocene resulted in the formation of the Farallon Negro volcanic center. Volcanism at Farallon Negro ini-

Jan 8, 1988