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By J. H. Brophy, S. J. Michalik

A constitution diagram for the system Mo-Ir has been determined. The maximum solubility of iridium in molybdenum is 16 at. pct at 2110ºC and decreases to less than 5 at. pct at 1500°C. The solubility of molybdenum in iridium is 22 at. pct. Three intermediate phases appear in the system: 8 MoJr, having the p-tungsten structure; a phase, a cornplex tetragonal structure; and the hcp ? phase. Metallography, melting point determinations, X-ray diffraction and fluorescence, and electron micro-probe unalyses were employed in establishing the diagram. PREVIOUS to the present investigation, the intermediate phases in the Mo-Ir system were identified, but no detailed account of the phase diagram has been reported in the literature. Raub1 investigated alloys of Mo-Ir over an extensive range of composition between the temperatures of 800º and 1600°C. The in-termetallic compound MosIr was found to exist with nearly pure molybdenum, as the solubility of iridium in molybdenum was not detectable parametrically in this temperature range. MO3Ir was found to be iso-morphic with a ß-tungsten type structure, having a parameter of 4.959Å. An intermediate hcp phase, designated as the ? phase, ranged in composition from 52 to 78.5 at. pct at 800ºC, and from 41 to 78.5 at. pct Ir at 1200°C. Parameters noted for the ? phase were as follows: at 42.7 at. pct Ir, a = 2.771i0, c = 4.4366, c/a = 1.601; at 78.5 at. pct Ir, a = 2.736A, c = 4.378A, c/a = 1.600. Molybdenum was found to be soluble in iridium up to 16.5 at. pct Mo (83.5 at. pct Irj, with the parameter of iridium increasing to 3.845A at the solubility limit. Knapton,2 who investigated alloys between 15 and 85 at. pct Ir, essentially agreed with Raub's data, but, in addition, found a a phase in as-melted alloys near 25 at. pcto Ir. The oaphase lattice parameters were a = 9.64Å, c = 4.96Å, c/a = 0.515. The a phase was replaced by the 8 -tungsten phase on annealing at 1600°C. Knapton concluded that the a was stable only at elevated temperatures, and placed the composition of the a phase at approximately 30 at. pct Ir. The intermetallic compound Mo3Ir, with a lattice parameter of 4.965A, was included among the 8-tungsten structures reported by ~eller.' Matthias and Corenzwit,4 and Bucke15 studied the superconducting nature of MosIr, and reported a superconducting transition temperature of 8.$K. The present investigation describes the phase relationships in the Mo-Ir alloy system determined by melting point measurements, X-ray diffraction and fluorescence, and metallography. EXPERIMENTAL PROCEDURES Alloys for the determination of the phase diagram were prepared from powders. Commercial 99.9 pct Mo from Fansteel Metallurgical Corp. and 99.9 pct Ir powder from J. Bishop and Co. Platinum Works were used. The powders were weighed to nominal compositions, mixed, and then pressed, without binder, into compacts weighing 4 to 6 g. These were presintered in uacuo between 1200' and 1400°C for 1 hr, to reduce the degree of spattering during subsequent arc-melting. The compacts were arc-melted in a nonconsumable tungsten electrode furnace six times on alternate sides on a water-cooled copper hearth in an atmosphere of zirconium-getter ed argon at 500 mm of mercury pressure. In almost all cases, this procedure yielded buttons of satisfactory homogeneity. The composition of all melted buttons were confirmed by X-ray fluorescent analysis using the experimentally determined ratio of the iridium La1 line intensity to that of the molybdenum Ka1 line as a function of composition. In this determination four alloys analyzed by wet chemical methods were used as standards. An uncertainty range of ±1 at. pct has been attributed to all indicated compositions. All heat treatments and solidus measurements were carried out in tantalum resistance heating elements in vacuum conditions of 10-4 to 10-5 mm of mercury. A detailed account of this procedure has been reported by Schwarzkopf and Brophy.8 In the heat treatment and solidus measurements of iridium-rich alloys (50 at. pct Ir or greater), a tungsten lining was inserted into the tantalum resistance heating element because of a eutectic reaction which occurs between iridium and tantalum at 1948ºc.7 Heat treatments and solidus measurements carried out at compositions less than 40 at. pct Ir both with and without tungsten linings within the resistance

Jan 1, 1963

By E. Oshima

The only one reverberatory furnace smelting in Japan has been alive in Onahama Smelter, Onahama Smelting and Refining Company. The fuel for the reverberatory furnaces had been changed from bunker C-oil to mainly coal and partially waste tires to reduce the energy cost. The steam, recovered by the waste heat boilers of the reverberatory furnaces, has been utilized fully for the power generation to save the purchase power. Copper concentrate has been smelted also by injection through tuyeres in the P.S. converters since 1983. The amount of concentrate, treated in the converters, was increased by the construction of small oxygen plant in December, 1990. In the paper, the current practice and improvements in the reverberatory and converter operations are described.

Jan 1, 1993

By H. W. Hitzrot

Introduction Mechanical classifiers consist of a settling tank with parallel sides and a sloping bottom equipped with a mechanism which continuously agitates the pulp and removes the settled solids. The functions served by mechanical classifiers include the following: (1) allow the particles larger than the desired size to settle in the tank and provide an overflow product with a minimum of oversize particles, (2) produce an overflow product of sufficiently high solids content to meet the requirements of the subsequent processing steps, (3) agitate the pulp and permit separation of the entrapped undersize particles so they can report in the overflow product, and (4) drain and remove the underflow solids from the settling pool. Various types of mechanisms have been used to agitate the pulp and remove the settled solids, which include: (I) endless belt and chain drags, (2) reciprocating rakes, (3) helical flights mounted on a rotating shaft which extends through the settling pool, (4) helical flights inside of a rotating drum which contains the classifier pool, (5) troughed conveyor belts forming the settling pool in the concave curve of the belt closely following the tail pulley, and (6) low-speed impellers agitating the underflow settled in the apex of cone-type mechanical classifiers.23 The principal use of mechanical classifiers has been in closed¬-circuit wet grinding; however, this application has been widely displaced over the past few decades by the hydrocyclone. The hydrocyclone installation, including its feed pump, offers lower capital costs and requires less floor space. Nominally it produces an overflow prod¬uct of higher solids content, compared with a mechanical classifier, for comparable classification size up to 400µm (35 mesh). However, these advantages of the hydrocyclone are being gained at the sacrifice of some of the advantages of the mechanical classifier. The mechanical classifier requires less power and has lower maintenance costs than the hydrocyclone and its feed pump. The underflow product from the mechanical classifier can have a higher solids content and contain less entrapped undersize particles than the comparable product from the hydrocyclone. Consequently, it is capable of operating at a higher classification efficiency. A reduced circulating load24, 25 results under these conditions in a closed-circuit grinding application. Other applications for mechanical classifiers include desliming, dewatering, and washing operations. Sand-slime separation may be achieved using one or more stages of mechanical classifiers. Typical applications would include washing and deslimining of concrete aggre¬gate or sand; glass sand; abrasives; oyster shells; phosphate rock; iron, nickel, and chromium ores; alumina; zeolites; solar salt; and precipitates from chemical processing. As an example for the latter, Bitzer26 recommended the use of mechanical classifiers for countercur¬rent cementation-in-pulp of copper from the leach liquor. Similarly, mechanical classifiers are used in uranium processing27 for separation of the barren sands from the pregnant leach solution. Depending on the particle size and the effectiveness of the sand¬slime separation, the sand product from mechanical classifiers may be drained to solids content approaching 84%. Thus, mechanical classifiers are frequently used as dewatering equipment for various applications, such as: 1) Make a product having suitable solids content for subsequent transport by belt conveyor. 2) Provide a drained product to reduce cost of drying. 3) Provide a drained, washed, heavy sand product for return to the circuit. 4) Recover fine values from dilute plant effluents from coal, sand, phosphate-rock, and iron-ore washing plants and from steel-mill flue¬dust scrubbers. 5) Drain residual reagents of a preceding process stage prior to conditioning with reagents of different characteristics, e.g., cationic¬anionic flotation of phosphate rock, iron ore, etc. Design and Operating Features The proper selection of mechanical classification equipment re¬quires that the properties of the solids and liquid are adequately de¬fined. The following descriptive information regarding the solids is desired: feed rate, chemical and physical composition, density, tem¬perature, size analysis, and desired separation size. For the liquid used, the following data would be needed: feed rate, density, viscosity, pH, and temperature. Settling Area. The pool area required to permit a particle larger than the separation size to settle depends on the density and shape of the particle and on the density and viscosity of the pulp within the classifier pool. These criteria of the solids and pulp determine the settling rate which, in turn, determines the settling area required. Determination of the settling rate by long-tube batch tests can be used to calculate the settling area. A theoretical approach for estimat¬ing the settling area requirements is discussed in Chapter 1, Classifica¬tion Theory. This procedure can be used when neither test data nor empirical data from similar operations are available. Data on settling-area requirements and settling rates have been compiled by the manufacturers of mechanical classifiers. The approxi¬mate settling area required for a particular separation size can be estimated by using the curve shown in Fig. 65. This curve applies for closed-circuit grinding applications. For dilute velocity classification in open circuit, desliming, the length of the overflow weir, W, multiplied by the weir's mean distance from the feed opening, E, is used as the design criterion, instead of the settling area, in sizing the classifier. The area requirements, W X E, for dilute velocity classification can be expressed in terms of volumetric flow, Q, and settling rate, Rs, as follows: W x E = Q/(18.06 Rs) where W and E are feet, Q is gallons per minute, and Rs is inches per second. Overflow Weir. The hydraulic head, and thus the velocity at which the overflow crests the overflow weir, is one of the factors controlling the particle size of classification. Consequently, the length of the overflow weir must be designed to provide the overflow velocity which allows the particles larger than the desired size of classification to settle in the working area of the pool. Obtaining the necessary

Jan 1, 1985

By Cristina Escudero Martinez

The purpose of the Basic Oxygen Steelmaking (BOS) is to refine the hot metal produced in the blast furnace into raw liquid steel, which may be subsequently refined in the secondary steelmaking shop. Appropriated fluxes for the BOF could require low melting point to reduce the energy consumption, to reduce the energy to fluidity the slag at lower temperature and to melt at lower temperature than steel. It is necessary a good Tumble test result and the fluxes do not generate fines which could obstruct the blowing plugs, so we would have a homogeneous flux. A homogeneous chemistry that allows to control the basicity (Si and MgO levels) and a high absorption characteristics to combine the flux with P, S, Si, etc. are necessary. PASEK Dunite is an ultramaphic rock exploited in the north of Spain with a basic chemical classification being olivine and serpentine its principal minerals. It is formed by orthopyroxene, clinopyroxene and olivine in variable proportion. Dunite is a very homogeneous and hard stone. The main characteristics are: High hot and cold resistance to mechanical stress (shatter / crushing / abrasion), softening and melting points of a flux, not a refractory material, large porosity which is increased more than 20% when dunite is heated at BOF process temperature, and high absorption level, and stable components with low melting points. PASEK Dunite is a good flux in BOF process and its use is recommended if the BOF process produces pig iron with low Si content, generates high levels of slag and it is necessary maintaining the basicity index and the MgO amount and with high value of hot metal ratio (high amount of phosphorus to remove).

Aug 16, 2017

By Peter N. Blakey

The Kidd Creek orebody is a massive, base metal deposit with widths up to 600 feet and a strike length of 2,200 feet. The orebody dips eastwards at 70° to 80° and is amenable to sub-level open stoping. Five distinct zones (Fig.1) based on structural wall rocks and ore type parameters occur within the orebody. These zones are: Zone 1 - Mainly copper stringer ore Zone 2 - Massive pyrite ore Zone 3 - Massive sphalerite and chalcopyrite Zones 4 and 5 - Are banded and variable in nature Compressive strengths vary from zone to zone with an average of 22,000 p.s.i. and highs of 40,000 p.s.i. The mine is developed from a ramp extending from surface to the 1600 foot level and a 24 foot diameter, 3050 foot vertical shaft with main levels at 400 foot intervals. Sub-levels at 100 foot vertical intervals are driven from the ramp. transverse and longitudinal pillars to support the crown pillar left under the open pit. The layout allows an opportunity to enlarge or reduce pillar sizes and also to introduce fill on any horizon should conditions warrant. It also offers a large number of operating stopes spread through the various ore zones giving good ore blending characteristics. All ore from the stopes is handled through drawpoints using Wagner ST8 scooptrams with an average one way haul length of 650 feet.

Jan 1, 1976

By Lucas Hidalgo Quintana

El presente texto describe la mineralización de una de las estructuras mineralizadas de la mina Quiruvilca denominada la veta Papelillo. Para ello, se describe la fisiografía glaciar, la estructura mineralizada en volcánicos de la formación Callipuy del Cretáceo Superior a Terciario Inferior (mayormente de composición andesítica) intruida por pequeños stoks riolíticos, andesíticos y dacíticos. Asimismo, abundan la esfarelita, rubia, galena, marmatita en menor proporción enargita, rodocrocita y cuarzo. La caolinización es la alteración hipógena importante que ubica los cuerpos mineralizados, con texturas mayormente de relleno. Como controles estructurales, está la presencia de cimoides según el rumbo y buzamiento de la veta, la falla de relleno mineralizado fue normal con mayor potencia en los sectores de mayor buzamiento.

Nov 1, 1981

By George Oprea, Tom Troczynski, Joe A. Rigby, Waiman Lo

The refractory lining of the tuyere line in Peirce Smith converters have usually much shorter life in service than the rest of the lining, above or below it. There are already recognized chemical, thermal and mechanical factors which contribute to the wear of the lining. Our study presents the experimental results on microstructural changes occurring on these tuyere line bricks in a nickelcopper converter. The experimental bricks, after use at the tuyere line, were investigated using SEM/EDS and XRD techniques, to identify the mineralogical changes due to interactions with the nickel-copper matte or fayalite type slag at the process temperatures. The microstructural changes were correlated with physical and mechanical properties of the brick before and after use and a wear mechanism was suggested. Although the matte does not theoretically react with the mineralogical components of the brick, a partial oxidation of the Ni-Cu sulphides was identified using a line analysis by SEM/EDS in the densified layer at the hot face, which allows for mineralogical changes at the matte-refractory interfaces, with a high probability of microcrack formation due to the thermal cycling during use. These microcracks were also identified to open through inter-granular fractures parallel to the hot face, which appeared to be the main cause of lost lining during use. The reactions between the fayalite slag and refractory brick could eventually play a role in chemical dissolution of the brick only in the superficial layer at the hot face, as no slag components were identified deep into the brick.

Jan 1, 2009

Discussion of the paper of Archer E. Wheeler and Milo W. Krejci, presented at the Butte meeting, August, 1913, and printed in Bulletin No. 80, August, 1913, pp. 1831 to 1880. BRADLEY STOUGHTON, New York, N. Y.:-I would like to ask Mr. Krejci whether they have tried mixing coke dust with the lime in those converters and, if so, if any advantage has been found. It has been a very great advantage in the steel converters where coke dust has been mixed with the lime near the tuyeres. Of course the agent that causes the corrosion at that point is oxide of iron, and coke seems to prevent the formation of the oxide of iron and so protect the tuyeres. It would be interesting to know if it had the same effect in the copper converters. MILO W. KREJCI:-We have never tried it here. At Anaconda the tuyere section was formerly put in by mixing magnesite with tar and ramming that into the tuyere section, but they have abandoned that. Might I ask how you would apply the coke dust? How get it into the tuyere section ? How hold it there ? BRADLEY STOUGHTON :-The coke dust is mixed with silica and clay and then rammed in and burned. I suppose in your converters it would be mixed with the magnetite or magnesite and then agglomerated with the material in place. MILO W. KREJCI:-We never tried it here. PROF. JOSEPH W. RICHARDS, South Bethlehem, Pa.:-I would like to ask how thick this deposit of magnetite is on the lining. How thick a layer would be practical ? MILO W. KREJCI:-About 21 in. would be sufficient. SOLOMON LE FEVRE, Mineville, N. Y.:-I was interested to hear of the use of magnetite in connection with the copper smelter. I am interested in the mining of magnetite, but was not aware that it had been used at all' by the copper metallurgists.

Jan 11, 1913

By Robert M. Grogan

Fluorspar is the commercial name for fluorite, which is the mineral having the composition CaF2, calcium fluoride. Its valuable properties are due to its content of fluorine, and it is the only important commercial source of that element. The name is derived from the Latin fluere, to flow, because the mineral melts easily and is a good flux. History The mineral has been known since Greek and Roman times, when it was used for drinking cups and ornamental slabs. Various peoples including the Chinese and the American Indians have carved ornaments from large crystals. Its usefulness as a flux was known to Agricola, who wrote about it in 1546.1 Mining began in England about 1775 15 and at various places in the United States during the period 1820 to 1840.28 Substantial production was achieved first in the period 1888 to 1900 in consequence of the development and widespread adoption of the basic open hearth method of making steel, in which fluorspar is used as a flux. After that, production and usage was stimulated by growth in the steel, aluminum, chemical, and ceramic industries and accelerated by World Wars I and II. The organic fluorides called "Freons" entered the picture in 1931, and the use of anhydrous HF as a catalyst in the manufacture of alkylate for high octane fuel blends began in 1942. The development of the froth flotation process in the late 1920's, a differential flotation scheme for separating fluorspar from galena, sphalerite, and other minerals in the 1930's, and the application of heavy-media concentrating methods to the treatment of low-grade ores in the 1940's were outstanding technological contributions that enabled increased production of fluorspar concentrates. Composition and Properties Fluorite in its purest form contains 51.1 pct calcium and 48.9 pct fluorine. Substitution of small percentages of cerium and yttrium for calcium have been noted.4 Mechanical inclusions of gases and fluids such as petroleum and water, and of other solid minerals such as pyrite, marcasite, and chalcopyrite are common. The fluorspar of commerce contains attached and admixed mineral impurities, chiefly calcite, quartz, barite, celestite, and various sulfides. It is interesting that the purest commercial form in which the mineral is sold, acid grade concentrate, contains a minimum of 97 pct CaF2, and as such is of comparable purity to many reagents used in chemical laboratories. Fluorite has a marked tendency to occur in large, well-formed crystals sometimes six to ten inches on a side. It crystallizes in the isometric system, frequently forming cubic and octahedral crystals or combinations thereof. The mineral also occurs in massive and in earthy forms, and in crusts or globular aggregates with radial fibrous texture. Crystalline fluorspar exhibits a great array of colors, from colorless and water clear to yellow, blue, purple, green, rose, red, bluish and purplish black, and brown. The colors are often arranged in alternating bands parallel to cube faces. They may be altered by exposure to X-rays, heat, ultraviolet light, and pressure, and are caused by a variety of factors including the presence of trace impurities and displaced ions in the lattice. Sometimes the changes induced by X-rays can be reversed by ultraviolet radiation. The mineral has a hardness of 4, and is the model of that hardness on the Mobs scale.

Jan 1, 1960

By Fernando Linares Quiroa

El diseño y operación de la mina es un proceso importante para la construcción de un modelo de cuerpo mineralizado. En este sentido, el presente texto describe la trascendencia del modelo geológico, construcción y uso de este para la extracción del mineral, donde este es una representación de la realidad que se genera a partir de la información predecible y proporciona a los geólogos e ingenieros de planeamiento los recursos más adecuados para una efectiva selección y extracción del mineral tanto física como económicamente. Asimismo estos bloques permiten la manipulación de la información contenida en estos a una escala local y describen contenido sobre las características tales como: leyes, dureza, peso específico, RQD, litología, mineralización, alteración, metalurgia, etc. Finalmente, los usos que se dan a los modelos de bloques se aplican al control del mineral, planeamiento a corto, mediano y largo plazo, el cáculo de los recursos geológicos y las reservas minerales, constituyéndose en la base para la optimización y del diseño del tajo.

Jul 31, 2003

By J. W. Paul

The material that composes the roof in coal mines differs in strength according to variations in its physical nature and characteristics. Some material such as limestone and massive sandstone, (free from slits) and carbonaceous shale or bone, will often remain in place in narrow work without the use of timber or other means of support, but in wide places the use of supports becomes a real necessity. Roof material composed of shale, fire clay or small laminations of sandstone, or shale or sandstone interspersed with laminations of thin coal, presents conditions that call for timber support in both narrow and wide work. However, in narrow work such unstable roof material may be taken down until the strata assume an arch that is self-supporting for overlying firm strata such as a roof coal, a touch weather-resisting, shale, sandstone, or limestone. The proper support of the roof in roadways used for haulage or travel adds to the economy of operation by preventing delays in transportation and increases the safety of those who travel these roadways.

Jan 1, 1930

By Isaac A. Zlochower, Michael J. Sapko, Ronald S. Conti

The Bureau of Mines has developed a system using a high-speed electropneumatic mechanism for the rapid (grab) sampling of dusts and gases during an explosion. The sampling system consists of an aluminum housing that incorporates two 30-cm3 preevacuated glass vials with rubber septums. Upon actuation, the sampling probe needle is driven through the septum with a pressurized air pulse, filling the tube with gas and dust from the mine explosion. After a predetermined time, the sampling probe needle is retracted by a second high-pressure air pulse to its normal (quiescent) state, allowing the sampling tube to reseal. The onset and the duration of sampling are independently variable and controlled by a time-delay relay package. This technique enables the monitoring of pyrolysis-charring in coal particles and the generation and combustion of the pyrolysis vapors in both large- and small- scale explosions. The following results were obtained from full-scale dust explosion tests at the Bureau of Mines Lake Lynn Laboratory: (1) Gas sampling of the leading edge of the flame front shows the large concentration changes, characteristic of the flame front; (2) gas samples taken entirely in the flame zone consist of pyrolysis and combustion products with very low residual oxygen; and (3) the particles collected in the flame zone show signs of extensive pyrolysis and charring.

Jan 1, 1988

By Fathi Habashi, Michel Pagé, Farouk T. Awadalla

"Chrysotile asbestos can be dyed with certain dyestuffs to form stable coloured products whose hemolytic and cytotoxicity are greatly reduced. Examples of these dyes are Thiazol Yellow G (a monazo dye) and Trypan Blue (a diazo dye). Other dye stuffs, for example, basic dyes with triphenyl methane structure, Brilliant Yellow (a diazo dye), and Methylene Blue (a thiazine dye) colour asbestos but do not reduce its toxicity. Similarly alizarin, 8-hydroxyquinoline, and representative members of the phthalein group Bromocresol Purple and Phenolphthalein.IntroductionThere is, at present, a controversy regarding the dangers of asbestos fibres, and attempts are being made to render it less toxic. One of these attempts consisted of treating chrysotile asbestos with gaseous phosphorus oxychloride, POCI], at ambient temperature and in absence of moisture. The product had to be heated at 300°C to 500°C to expel exess entrained reagent to get a material containing 0.5070 to 5% phosphate groups and known as Chrysophosphate (Lalancette and Dunnigan, 1982). This product showed decreased toxicity as measured by its hemolytic effect on red blood cells (fable 1). In another approach, Flowers (1983) showed that when chrysotile asbestos was thoroughly mixed with a solution of ferric ammonium sulphate and the slurry was then filtered and dried, the product showed a marked decrease in cytotoxicity.In the present investigation the problem of toxicity was considered from a different point-of-view. Chrysotile asbestos is a hydrated magnesium silicate that can be represented by the structural model shown in Figure 1 from which it can be seen that magnesium hydroxide groups, Mg(OH)2, form a part of the silicate structure. Certain organic reagents that are known to react with magnesium ion in solution to form chelates were reacted with asbestos to find out whether a chelate was also formed on the fibres, and if so , what will be the hemolytic action and cytotoxicity of the product."

Jan 1, 1991

By John T. Bartlett

During the last decade, the increasing price of gold has renewed interest in the various metallurgical techniques for recovering precious metals. Much of the recent technology advance has been aimed at low grade deposits resulting in more complex processes and the necessity of tighter economic control. These factors along with more stringent environmental constraints have made it necessary to optimize each process to minimize energy and raw material consumption and reduce waste discharge. In most cases, detailed process analysis is required to detect inefficiencies and provide alternative solutions for process improvement. Also during this same period, the computer has become widely accepted and computerized mathematical modeling has been applied to an increasingly broader range of problems. Microcomputer based process modeling and simulation systems are routinely used to analyze and improve communition, hydrometallurgical, and pyrometallurgical processes for the recovery of precious metals.

Jan 1, 1987

By Sam Tour

SOME years ago, the writer described heat checks or thermal. cracks that occur in die-casting dies.1 The life of dies was considered in relation to the casting temperature, the material used for the die, and the heat to be absorbed by the die per unit volume of casting made. The mech-anism of the action that results in thermal failure of dies was described. In the present paper the writer proposes to describe some additional observations in regard to heat checking, make one or two suggestions as to the method or methods for further attack of the problem, and give the results obtained on some work carried out along these lines.

Jan 1, 1934

By Finlay, J. S.

THE North Lily Mine started its career in an unusual way it was discovered by a geologist. The remarkable circumstance of driving a 2400-ft. drift into an unexplored country and "hitting her on the nose," indicates that geologists are sometimes numbered among those whom the Lord loves. Diamond- drilling from surface had previously shown the presence and position of the Ophir beds, and surface showings of pebble-dykes, faults, and bleached rhyolite indicated conditions similar to those at the Tintic Standard "Pot-hole." The discovery drift from the Tintic Stand- ard 700-level reached the North Lily orebody in January, 1926, and development and stoping were continued by Tintic Standard men and equipment until the North Lily shaft was sunk and put into service in May, 1928. The discovery drift was run on a cooperative basis by the Tintic Standard and North Lily mining com- panies. Finding the North Lily mine cost the North Lily company $26,000. The Tintic Standard Mining Co. furnished men and equipment to drive this discovery drift at a fraction of what it would have cost to find the orebody from surface. This was the beginning of the friendly cooperation that still exists. Through Tintic Standard ground the Tintic Standard paid two- thirds of the cost and the North Lily one-third; the cost of the drift through Eureka Lily, into North Lily ground, was paid by North Lily. Later, another drift from the Tintic Standard 900-level was run on the same plan, excepting that Eureka Lily paid its share of the cost; that is, the Eureka Lily paid one-third of the expense in Tintic Standard ground, and the Eureka Lily two-thirds and North Lily one-third in Eureka Lily ground. This 900-drift was driven by North Lily from the Eureka Lily line to near the North Lily west side-line. From this point the Chief Consolidated continued the drift well into Tintic Bullion ground.

Jan 1, 1929

By C. P. Melloy

INTRODUCTION Ventilation practice at Isa Mine has been evolving since underground mining began in 1931. Much fundamental work has been conducted and reported regarding the primary ventilation system [Mathews K.E. et a1 (1967) and Brown J.R., Fisher T.H., Gibson K.L. (1979) 1. However, ventilation networks for subsystems such as open stopes, have evolved with experience. This paper outlines current ventilation practice for open stoping. Mount Isa Mine The mine is located in semi-arid North-West Queensland, Australia, and currently produces 20 000 tonnes of copper ore and 10 000 tonnes of silver-lead-zinc ore daily in each five day working week. Mining extends over a north-south strike length of 3.5 km with a maximum width across dip approaching 1 km, and is to a depth of 960 m (Figure 1). At present approximately 90% of the mine's copper production is from the 1100 Orebody, which contains ore reserves of 121 million tonnes of 3.2% copper. Open stoping is currently the mining method employed, and is used for primary stoping and pillar recovery adjacent to cement fill walls. The development of stoping methods in the 1100 Orebody has been documented by Kirkby R.W., Blair J.R., and Hutton R.C. (19781, Hornsby B. and Sullivan B.J. (1978), and Alexander E.G. and Fabjanczyk M.W (1981). As many as thirteen open stopes may be producing concurrently in the Orebody (Figure 2). The multiple silver-lead-zinc orebodies containing reserves of 54 million tonnes of 150 g/t silver, 6.4% lead and 6.5% zinc, are mined by open stoping, and cut and fill techniques. These methods have been documented by Goddard I.A. and Bridges M.C. (1977 1, and Goddard I. A. (1981 1. Only the ventilation of the 1100 Orebody open stopes will be discussed. However, the same ventilation philosophy is employed in open stope design in the lead orebodies.

Jan 1, 1981

By Simón D. Strauss

El presente estudio revisa la evolución de la producción de los principales minerales (plomo, cobre y plata) en Estados Unidos desde la década del 70, asi como su repercusión en Chile, Perú y algunos países latinoamericanos. En esta perspectiva, se pronostican algunos datos sobre los cambios políticos, sociales y económicos que ocurrirán dentro de los próximos 28 años y que afectarán la demanda de productos minerales y la capacidad para producirlos.

Nov 20, 1972

By J. O?Hanlon

The 4500 m3 blast furnace at the Redcar Works of ESC operates or, a burden of sinter and pellets, both materials being produced within the works. The early history of the furnace was given to the AIME Ironmaking Conference in 1981(1) when the burden was 65% superfluxed minter and 33% acid pellets. While ironmaking performance was adequate, it was felt that there were advantages to be gained by the use of fluxed pellets. For example, Kobe Steel had reported (2) that the rise of 30-50% dolomite fluxed pellets had enabled them to achieve better furnace permeability, higher gas utilisation, and enhanced ore/coke ratios. It appeared probable to BSC that the use of magnesia-bearing fluxed pellets would give a burden which had compatibility of softening and melting properties, i.e., fluxed sinter and dolomite pellets would result in a thinner cohesive zone because of closer matching of softening and melting temperatures, with consequential benefits on furnace stability and efficiency.

Jan 1, 1983

By P. Hajiani

"It has been shown that the economics of the hydrometallurgy of a rare earth process depend upon the amount of acid needed to extract the metal per unit mass of the ore. In this paper, we show how to save a significant amount of acid treating a low grade ferrocarbonatite and extracting +99% constituent REEs. In order to save the acid, the alkali earth metals, originated from gangue minerals were removed prior to REE leaching. Moreover, the operating conditions of leaching reactor are optimized to leave iron behind and to leach REE selectively and effectively. Consequently, the hydrochloric acid consumption per ton of ore drops significantly compared to other processes.INTRODUCTIONThe viability of an extractive metallurgy process of rare earth elements (REEs) to produce an intermediate mixed concentrate depends upon a low-cost simple flowsheet (Bogner, 2014). It has been known that regardless of the mineral, the acid consumption is intimately associated with the operational cost (OPEX) of any REE project, particularly when the initial REE grade is low. Besides acid, the other reagents consumed in the process impact the OPEX. This makes minimizing acid consumption in an REE method a critical priority. The acid consumption in different projects has been reported to be equivalent of 100 kg to 1000 kg HCl (100%) per tonne of ore (Average: 500 kg/t HCl 100%) (Bogner, 2014). This value is strongly correlated with the grade of REE in the ore and therefore in the pre-concentrate.Previously, it has been stated that the Montviel flotation concentrate requires about 300 kg HCl (100%)/t ore to attain +99% REE recoveries using direct leaching (Yu, Verbaan, Pearse and Britt 2013). Hence, for an ore feed rate of 2500 t/d ROM, a large acid production plant (744 t/d HCl 100%) would be required for acid leaching of such concentrate. The equivalent Chlor-alkali plant is estimated to cost M$ 150 (CAPEX) and may require 100 MW power. Therefore, there is a necessity to improve the extractive hydrometallurgy and to reduce the acid consumption significantly."

Jan 1, 2016