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By George G. Griswold

The Parkes process of lead refining is based on the fact that when zinc isemixed with molten lead bullion, it forms an alloy with the silver and gold. This alloy freezes at a higher temperature and is of a lower specific gravity than the lead. With proper regulation of the temperature, the alloy solidifies and floats on top of the lead bath, from which it is removed and the silver, gold, and zinc recovered. The alloy should contain the maximum amount of silver (Ag2Zn,). The usual practice is to make two additions of the zinc. The first addition gives an alloy, or zinc crust, the zinc of which is saturated and is treated for the recovery of its zinc and silver; the crust from the second addition contains an excess of zinc and is used for the first addition on the following kettle. The crust produced depends on a number of items: purity of bullion, temperature, care with which kettle work is performed, addition of zinc, pressure used in expelling lead from crust, etc. It is desirable to make the concentration of silver in the zinc crust as high as possible; otherwise an undue amount of byproducts is produced, which greatly increases the cost of refining. It was formerly considered good practice to make a concentration of 20 :1. That is to say, if the lead bullion assayed 150 oz. dorê per ton, the assay of the retort bullion recovered from the zinc crust should be 3000 oz. per ton. As a matter of fact, the concentration should never, except on bullion assaying over 300 oz., be so low, and with low-grade bullion it should be 50:1. In one lead refinery where, ordinarily, the work was good, on a bullion averaging about 200 oz. per ton, a concentration of 24:1 was common; but how it was obtained was known only to the foremen. Changes occurred, and under new foremen the kettle work fell off badly; the problem then was to bring about this greater concentration. As is customary, the refincry had a "schedule" for its 60-ton kettles, which gave the number of pounds of zinc required for all possible assays, on both the first and the second zincing. Using this schedule, high temperatures

Jan 1, 1924

Jan 1, 1912

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

Upon the recommendation of its Technical Committee, The Com-pressed Air Society has adopted the following definitions of, certain terms. Displacement.-The displacement of an air compressor is the volume displaced by the net area of the compressor piston. Capacity.-The capacity should be expressed in cubic feet per minute, and is the actual amount of air compressed and delivered, expressed in terms of free air at intake temperature and at the pressure of dry air at the suction. Volumetric Efficiency.-Volumetric efficiency is the ratio of the capacity to the displacement of the compressor, all as defined above. Compression Efficiency.-Compression efficiency is the ratio of the work required to compress isothermally all the air delivered by an air compressor to the work actually clone within the compressor cylinder, as shown by indicator cards, and may he expressed is the product of the volumetric efficiency (the intake pressure and the hyperbolic logarithm of the ratio of compression), all divided by the indicated mean effective pressure within the air cylinder or cylinders.

Jan 8, 1918

By Stuart S. Forbes

Changes and additions made to the Canadian Copper Refiners Ltd. electrolytic refinery between 1949 and 1955 are reviewed. The effect of high current density on current efficiency and section work is discussed. OPERATING and physical changes which have been made to the electrolytic tank house of Canadian Copper Refiners Ltd. suggest a review of present conditions. Previous papers1 a have described the general arrangement. Since then, expansion, closer anode spacing, and higher current densities have increased annual cathode capacity from 112,000 to 182,000 tons. A recent extension to the tank house, begun in the spring of 1954 and completed in August 1955, was undertaken to handle the monthly production of about 3,000 tons of anodes from Gasp6 Copper Mines. The anticipation of these additional receipts and large stocks of unrefined copper caused by the increase in custom smelting at Noranda made it necessary to increase production without immediately available additional cell capacity. Accordingly, from December 1954 through March 1955, the tank house was operated successfully at a current density of 24.0 amp per sq ft. Again, as of September 1955, one of the two tank house circuits is being operated at this high density. Although other refineries may operate at current densities higher than 24 amp per sq ft, none is doing so with anodes which contain silver and gold content as high as those being refined at Canadian Copper Refiners Ltd. General Arrangement The original building consisted of two parallel bays, each 660 ft long and 60 ft wide, with a 24 ft pump bay extending along the west side of the building. A previous expansion, completed in 1939, was to the west of the original building and consisted of two 60 ft parallel bays 240 ft long. The length of these two bays was extended in 1954 to 540 ft, adding one stripper and 12 commercial sections to No. 2 electrical circuit. Space is still available for an additional six commercial sections and one stripper section. Construction on these seven sections started in September 1955, and will be completed early in 1956. The tank house capacity will then be 206,000 tons of cathodes per year. In the two tank houses there is a total of 900 cells distributed as follows—commercial, 468; stripper, 36; total, 504 in the No. 1 tank house (No. 1 circuit): commercial, 360; stripper, 36; total, 396 in the No. 2 tank house (No. 2 circuit). Section arrangement consists of nine cells to a tier and two tiers or 18 cells to a section. Anodes, weighing 620 lb each, are cast with a Baltimore groove and are refined by the multiple system. Solution enters each cell through a bottom inlet at one end and overflows at the top of the opposite end. Rate of solution flow through the cells is controlled by valves at each section to 4.5 gpm. Electrical power to the tank house is now provided by nine 675 kw motor generator sets. Each set consists of a 135-12 v 5000 amp dc generator, driven by a 980 hp 2300 v synchronous motor. When No. 1 circuit is operated at 22,000 amp, five sets, at 4400 amp each, are connected in parallel. As the maximum allowable current on No. 2 electrical circuit has recently been set at 18,000 amp, only four sets are connected in parallel on No. 2 circuit. A tenth set, now being installed, will be reserved as a spare, for use in either circuit as required. Recent Changes In 1949, the annual capacity of the tank house was increased from 112,000 to 132,000 tons by closing the

Jan 1, 1957

By Young C. Kim, William C. Cassun

A proper economic analysis of pit slope design must reflect the trade-off between the benefits and the increased risk of slope failure inherent in steeper slope angles. This paper describes the results of a case study in which a newly developed probabilistic economic analysis procedure was applied to the selection of a pit slope design at a large porphyry copper mine in the southwestern U.S. The paper also demonstrates the availability of the required data such as the probability of failure schedule and cost of failure models, under "real-life" conditions.

Jan 1, 1978

THE CHAIRMAN (EDWIN LUDLOW, Lansford; Pa.).-Mr. Wang's paper reminded me that some years ago I was asked if I would not go into the Tien Tsin district and take charge of Americanizing a Chinese mine. They said the trouble was that, while wages were only 9 c. a day, the coke cost $2 a ton delivered to the cars. Reading Mr: Wang's paper would indicate the need of a little Americanization in the Manchurian field. Wages are only 40 c: a day, yet the cost is $3.60 a ton. An efficiency engineer would have an excellent opportunity in that field. T. T. READ, New York, N. Y.-It is always true that low-priced labor is inefficient, wherever we find it; for example, in Japan, where wages are low, but not so low as in Manchuria, working costs are not relatively much lower than ours. Mr. Padshah, describing his experience in India, once told me that, while wages were only about one-fifth as high

Jan 4, 1918

By George F. Comstock

It has been reported recently1 that the hardenability and toughness of forging steels may be improved appreciably by the use of complex deoxidizers containing titanium, aluminum, and vanadium. In the light of these results, an investigation of similar deoxidizers containing various hardening and grain-refining elements was proposed. These deoxidizers included three manganese-silicon-aluminum-titanium al- loys with and without boron and calcium, three titanium-aluminum ferroalloys with vanadium, molybdenum and zirconium, respectively, an aluminum-zirconium ferroalloy, and plain ferroboron. Their approximate compositions are given in Table I, and the results obtained so far in the investigation of their effects on steel are presented in the later parts of this paper. Method of Using Deoxidizers The deoxidizers listed in Table I were used in three series of 0.40 per cent carbon steels, which were melted in a small basic-lined induction furnace and cast in the form of I7-lb. ingots about 2.75 in. square. The basic charge for each heat was 10 lb. Armco iron and 6 lb. of clean low-carbon sheet-steel scrap. When melted, this charge was deoxidized with 15 grams of ferrosilicon, then about 500 grams of pig iron was added to give the desired carbon content. After the pig iron was dissolved, nickel shot and high-carbon ferrochromium were added if required, and additions of 80 per cent high-carbon ferromanganese and 50 per cent ferrosilicon were made to provide the proper amounts of those elements. Aluminum was added to every heat 2 min. after the silicon, for grain-size control and preliminary deoxidation. One minute later the special deoxidizer was added to the clean bare surface of the molten steel, and 2 min. afterward the steel was poured into the ingot mold, using a hot-top and pipe-preventing compound after pouring to decrease the depth of the shrinkage cavity. The special deoxidizers were used mostly in the form of about ½ to¾-in. lumps,

Jan 1, 1942

By George F. Comstock

It has been reported recently1 that the hardenability and toughness of forging steels may be improved appreciably by the use of complex deoxidizers containing titanium, aluminum, and vanadium. In the light of these results, an investigation of similar deoxidizers containing various hardening and grain-refining elements was proposed. These deoxidizers included three manganese-silicon-aluminum-titanium al- loys with and without boron and calcium, three titanium-aluminum ferroalloys with vanadium, molybdenum and zirconium, respectively, an aluminum-zirconium ferroalloy, and plain ferroboron. Their approximate compositions are given in Table I, and the results obtained so far in the investigation of their effects on steel are presented in the later parts of this paper. Method of Using Deoxidizers The deoxidizers listed in Table I were used in three series of 0.40 per cent carbon steels, which were melted in a small basic-lined induction furnace and cast in the form of I7-lb. ingots about 2.75 in. square. The basic charge for each heat was 10 lb. Armco iron and 6 lb. of clean low-carbon sheet-steel scrap. When melted, this charge was deoxidized with 15 grams of ferrosilicon, then about 500 grams of pig iron was added to give the desired carbon content. After the pig iron was dissolved, nickel shot and high-carbon ferrochromium were added if required, and additions of 80 per cent high-carbon ferromanganese and 50 per cent ferrosilicon were made to provide the proper amounts of those elements. Aluminum was added to every heat 2 min. after the silicon, for grain-size control and preliminary deoxidation. One minute later the special deoxidizer was added to the clean bare surface of the molten steel, and 2 min. afterward the steel was poured into the ingot mold, using a hot-top and pipe-preventing compound after pouring to decrease the depth of the shrinkage cavity. The special deoxidizers were used mostly in the form of about ½ to¾-in. lumps,

Jan 1, 1942

By R. G. Hudson, L. Yang

In electrochemical separation of metals, it is necessary to control the potential applied between the electrodes so that only the desired electrode reactions can occur. A knowledge of the minimum potential needed for a given electrode reaction to take place continuously is therefore pertinent to the success of the process. In molten chloride systems at high temperatures, activation polarization is believed to be small,1,2 so the minimum potential needed for the electrolytic decomposition of a molten metal chloride MC1n (n = valence of the metal ion) is therefore essentially the equilibrium electrode potential E of the metal-chlorine galvanic cell in the solution. If the decomposition products are pure M and chlorine at a pressure of p atm, E*Eo-rt/nf in [i] Since EO (the standard electrode potential of MCln) is a constant at a given temperature and R (the gas constant), and F (the Faraday's constant) are all constants, the variation of E with the electrolytic bath conditions is largely determined by the effect of the latter on the activity of the metal chloride amcln (standard state = pure MC1,). Measurement of E and study of how amCin varies with temperature and concentration, the nature of the solvent, and that of the coexisting solutes are therefore pertinent in the selection of optimum operating conditions for the electrochemical separation of metals. Because of its low melting point, high decomposition potentials of its components, and the low solubility of heavier metals in it, the LiC1-KC1 eutectic melt is a useful solvent and has been used in the electrodeposition of Mo,3Ce,4 and u5 metals. Senderoff and Brenner3 measured the potentials of Zn/Zn++, Fe/Fe++, Cu/Cu+, Mo/Mo+++, and Ag/Ag+ against an Ag/AgCl (pure) reference electrode at 600°C and a concentration of 4.1 mole pct of each chloride and came to the conclusion that complex ion formation occurred in some of the melts. Laitinen and LiUe gave an electromotive force series at 450°C for a number of {Metal/Metal ions (unit activity)) electrodes in LiC1-KC1 eutectic melt, the {pt/pt++ (unit activity)) electrode being chosen as the reference. Walker and Danly7 studied the ther-modynamic properties of NiC12 in LiC1-KC1 eutectic melt over a wide range of temperatures and concentrations by measuring the electrode potential between a nickel electrode and a chlorine electrode in the melt. Since potential measurements made by using the chlorine reference electrode are of more direct significance both in the study of electrochemical separation and in the evaluation of the thermodynamic properties of the melt, it is therefore decided to measure the electrode potentials of cells of the type {M/MCin (mole fraction = N) in LiC1-KC1 eutectic melt/Cl2 over a range of N values and temperature for a number of metal chlorides. The results should indicate the order by which the various metals plate out on the cathode and how this order is affected by temperature and the concentration of the metal chloride in the solu-

Jan 1, 1960

By Theodore P. Holt

The Mines Operating Co.'s plant at Park City, Utah, was designed to treat the low-grade fillings in the old stopes of the Ontario mine. These fillings carry 6 to 14 oz. of silver, 1 to 2 lb. of copper, 0.01 to 0.015 oz. of gold, and a small percentage of lead and zinc. The treatment consists in mixing the crushed ore with coal dust and salt and then roasting in a new type of furnace by combustion of the contained fuel. The roasted ore is leached with an acid salt solution to dissolve the silver, gold, copper, and lead. At present these metals are precipitated together on scrap iron and the product sold to a refinery. The only new feature of importance in this scheme of treatment is the roasting process, which makes possible the chloridizing of ores without any loss of the valuable metals, and at a very low cost. History of the Development of the Process The roasting process- had its beginning in connection with research work at the Utah State School of Mines. N. C. Christensen, Jr., who held one of the School of Mines' fellowships, was engaged in some experimental work on blast roasting. At the same time I was doing some work on the insoluble gold in Mercur base ores. Mr. Christensen roasted some of this base gold ore in his pot furnace, and found that when he mixed a sufficiently low percentage of fuel with the ore, it did not sinter but roasted to a leachable product. We considered this a new application of blast roasting, and proceeded to test out a large number of ores in several different types of small roasters which we constructed.. This experimental work continued well into the summer of 1911. In August, 1911, the Consolidated Mercur Gold Mines Co. installed a small roaster of our design, which we proposed to operate on the counter-current principle, feeding mixed ore continuously on top, and at the same time drawing off the roasted ore from the bottom. By proper adjustment we hoped to maintain a permanent roasting zone near the center of the column, the ore moving downward as roasted. However, we

Jan 1, 1915

By C. P. Proctor, J. T. Crawford, John Griffen

TROUGH washers were among the earliest methods used for concentrating ores; they are referred to by Agricola about the middle of the sixteenth century as already being used while the hand- operated jigging sieve had only recently come into use. For cleaning coal the trough washer apparently was first used in France and Belgium about 184.1. The early trough washers were intermittent and the refuse at least was shoveled out by hand. The invention of a continuous jig washer, by a Frenchman, Berard, in 1848, acted as a stimulus to coal washings in jigs and retarded progress in trough washing in the nineteenth century. Chapman and Mott 1 describe a number of trough washers developed in England and on the Continent from 1850 to 1900, such as the Bell and Ramsay, Ramsay, Bell and Green, Wunderlich, Scaife, McLellan, Lodge, Elliott, Blackett, and Murton. The later of these removed the refuse mechanically, either intermittently or continuously, but in ways that interfered with stratification. THE RHEOLAVEUR A Belgian engineer, Antoine France, is responsible for the development of means to remove refuse materials continuously from a trough washer, while preserving stratification and preventing the loss of coal with the refuse. The methods and equipment devised, which have become known throughout the world as rheolaveur, were initiated at the St. Nicholas mine near Liége, Belgium, about 1912. This installation consisted of two units to wash 20 to 8-mill. (3/16 to 5/16-in.) and 8-mm. to 0 (5/16-in. to 0), each having only two superimposed inclined troughs. The first few "Rheo" boxes on the top trough removed refuse directly and the remaining Rheo boxes discharged impure refuse to the lower trough for rewashing. On the lower trough the first few boxes removed refuse and the following boxes removed a middling product used for boiler fuel or other purposes. The overflow of both troughs was joined to form the washed coal product.

Jan 1, 1950

By AIME AIME

BY COMBINING the sessions on reduction and refining of copper, lead and zinc it was possible to devote an entire day to nonferrous metallurgy. Four interesting papers were presented at the morning session and four more in the afternoon. In the discussion of papers on copper, F. R. Pyne acted as chairman, with H. M. Shepard as vice-chairman, whereas for those on lead and zinc, F. F. Colcord acted as chairman, with C. R. Hayward as vice-chairman. The paper by P. D. I. Honeyman on the smelting of raw concentrates in the reverberatory furnace was presented by Mr. Pyne and discussed by Messrs. Mazany and Hayward. The interesting paper by E. S. Bardwell, com- paring the use of various fuels in copper-refining furnaces at Great Fails, Mont., was presented by Mr. Shepard and discussed by Messrs. Boggs, Gronningsater and Hayward. That by R. G. Knickerhocker, describing the stationary basic copper converter at Messina, Transvaal, was presented by Mr. Pyne, together with a written discussion by c. R. Kuzell, of the United Verde Copper co. It was an interesting illustration of how well-known processes may sometimes be modified to fit into unusual conditions.

Jan 1, 1932

By Alfred L. Beebe

to. A strip of thin platinum foil, about one inch square, is perforated with pin-holes over its entire surface. It is then bent into the shape of a letter U, and its opposite corners are connected by short pieces of platinum wire, so as to form a sort of basket. In this is placed a piece of amalgamated zinc of convenient size, and the whole is suspended by a stout platinum or silver wire in the bottle containing the solution to be reduced. Another strip of platinum foil,

Jan 1, 1886

By William A. Kapp

INTRODUCTION Coal is being mined from beneath residential areas, structures, bodies of water and other surface features in the coalfields to the north, south and west of Sydney. The particular problems faced by mine operators in these areas vary considerably due to differences in the overlying strata, the variation in the depths of cover and also depend on the number of seams being mined. Detailed subsidence work first commenced in the Southern Coalfield in 1965 and is now being carried out over areas of extraction at roost collieries. The analysis of the results of the early investigations and of the work which continues in other areas has shown that there is a consistent relationship between subsidence and mine geometry and has led to a reliable empirical method for the prediction of subsidence. In addition, particular aspects of each of the studies in the Southern Coalfield results in a clearer understanding of strata movements and of the resulting subsidence. The features of a subsidence trough apply generally to all areas but the magnitudes of specific features vary according to the stratigraphy of the particular coalfield. The aim of the subsidence work is to quantify the effects of subsidence for a range of mining geometries and mining conditions to enable the maximum safe recovery of coal from beneath surface features. The importance of local subsidence investigations is becoming more evident to mine operators and to authorities or organisations with surface interests. The subsidence work also provides important information on the stabilities of pillars of coal which remain unmined between panels of extracted coal. These pillars are not extracted either because of poor mining or geological conditions, or because pillar extraction is not part of the particular mining operation. Subsidence studies over these coal pillars clearly establish whether the pillars have remained stable or have failed to support the overlying strata. With subsidence studies continuing over several years, it is possible to assess the stabilities of these pillars on a long term basis. BACKGROUND TO THE STUDY OF SUBSIDENCE Geographical setting Most of the black coal production in Australia comes from the Sydney Basin. The coal seams extend for approximately 350 km along the coast of New South Wales and inland for distances up to 150 km. The City of Sydney is located near the centre of the coastal extent of the Basin where coal has been mined at a depth of 900 m. The Sydney Basin is part of the Main Coal Province of NSW and is divided into several coal- fields. The Southern Coalfield to the south of Sydney contained 15 operating mines and produced 12.7 million tonnes of raw coal during the 12 months to June 1981. The collieries discussed later are shown in Fig. 1. The prominent topographical feature of the area is the Illawarra Escarpment which rises to 400 m above sea level, or 300 m above the coastal strip along the South Pacific Ocean. The escarpment is mainly sand- stone and the weathering of the cliff line has resulted in a covering of talus material at its base. Several collieries are located near the seams which outcrop along the escarpment. The city of Wollongong is located in a scenically attractive area on the coastal plain. The suburbs of Wollongong extend north along the coastline, south to beyond Lake Illawarra and west to the lower slopes of the escarpment. The Illawarra Escarpment forms the eastern boundary of the Woronora Plateau. On a regional scale the surface dips gently to the west and thus forms a watershed for the rivers, most of which flow in a general north westerly direction, sometimes forming steep gorges in the sandstone. These rivers join the Nepean and Hawkesbury River system and flow into the Pacific Ocean north of Sydney. Seven dam have been constructed over the Southern Coalfield (Fig. 1) and with one large dam further to the west, their stored waters provide the needs of the Cities of Sydney and Wollongong and the surrounding districts. A large part of the area affected by mining is the undeveloped bushland of the associated catchment areas. In general no special precautions have been required with respect to subsidence with the exception of the dam structures and stored waters. With the increase in coal mining activities and the expanding residential development south of the City of Campbelltown in the outer Sydney Metropolitan area, subsidence is becoming an increasingly important area of research. Structures which have been affected or considered are townships and extensive residential areas, buildings of historical importance, major tollways, and a high pressure natural gas pipeline. The subsidence effects of mining beneath natural features within national parks is coming under study as mining approaches these areas. Geological setting The coal seams of the Southern Coalfield lie within the Illawarra Coal Measures. They contain high rank coking coal used in the local steel industry and for export. The Bulli Seam is mined extensively through- out the Southern Coalfield with the lower Wongawilli Seam being second in importance with regard to coal production. The top of the Bulli Seam is taken to be the marker horizon between the Permian Coal Measures

Jan 1, 1982

By K. H. Heino

Two 200 T/da Lurgi Turbulent Layer fluid-bed roasters have been in operation at Canadian Electrolytic Zinc Limited since 1966. Zinc concentrate containing 52-54% Zn, 9-11% Fe,31-33% S is treated to produce calcine containing 60-62% Zn,10-12% Fe, .l-.3% S as sulphide and .8-1.2% S as sulphate. Each roaster train comprises a waste heat boiler, two cyclones in parallel and a two-field electrostatic precipitator. Distribution of calcine collected is 10-15 percent from bed overflow, 55-65 percent from boilers, 20-25 percent from cyclones and 2-6 percent from precipitators. A Monsanto designed contact plant produces acid at a rated capacity of 365 T/da (100% H2304 basis) from roaster gases diluted to 7.0-7.5% S02. The process steps are gas cleaning and cooling by weak acid scrubbing, acid mist removal in electrostatic precipitators, gas drying in 93.5-96.5% H2304, conversion of sulphur dioxide to sulphur trioxide in a four-pass converter containing vanadium pentoxide catalyst, and absorption of the sulphur trioxide in 98.1-98.3% H2S04. Product acid is 93.2% H2S04. Conversion efficiency is 97-98 percent and overall sulphur recovery from concentrate is 87-90 percent. The adjacent fertilizer plant has a capacity of 135 T/da of P205. The products are 54 percent phosphoric acid, run-of-pile triple superphosphate, and diammonium phosphate. The plant is of Ugine-Kuhlman design and consists of an air-cooled multiple tank reactor, a tilting pan filter, and a forced circultation vacuum evaporator. Triple superphosphate is made on a continuous belt and partially cured in a rotary dryer. Diammonium phosphate is produced by ammoniating phosphoric acid in a single stage reactor. The slurry of ammonium phosphate is granulated in a blunger and dried.

Jan 1, 1970

By H. J. Rahilly

MINE FIRES, in the Butte District, have been a source of trouble and expense for the past thirty years, for while the actual fire area in most of the mines has been comparatively small, the handling of the gas has generally been a much larger problem than the actual putting out of the fire. The chief causes of these fires have been: Defective electrical equipment, incendiarism or carelessness, and spontaneous combustion. While fires from the first two causes have been numerous, they have generally been discovered and extinguished before they became very serious. Spontaneous combustion has been the cause of most of the large fires. Sufficient heat has been generated by ground movement, high rock temperatures, and decomposition to set fire to inflammable materials (tarred ropes, canvas, dry timber, manure, and hay) left in the gobs some years ago. Also, the heat resulting from the oxidation of the fine broken sulfide ore is sometimes sufficient to set fire to the timbers. In 1906, a fire was discovered on the 1100-ft. level of the Minnie Healy mine about 400 ft. (122 m.) west of the Minnie Healy shaft, which presumably originated, in one of the upper levels, from spontaneous combustion. The fire was walled off by concrete bulkheads and a pillar of ground 100 ft. (30 m.) thick was left below and mining resumed underneath. During subsequent operations, this pillar became cracked and broken and the fire passed into the new workings. It travelled downward and spread rapidly until, in 1917, it had become necessary to wall off portions of the Tramway, Leonard, and West Colusa mines on the 1300, 1400, 1500, 1600, 1800, 1900, and 2000-ft. levels, and the top of the 2200-ft. raises were sealed. The fire area now consists roughly of a block 1000 by 1200 ft. and from the 600-ft. to the 2200-ft. level. This block is composed largely of a series of parallel northwest-southeast nearly vertical veins and stringers that form a large orebody, with no well-defined walls but badly fractured and faulted, and known locally as "Horsetail Ore." These veins, or orebodies, are mined by the square-set or the rill method.

Jan 2, 1922

By J. L. Rodda

THE desirability of a uniformly sized abrasive for metallographic polishing has probably been recognized in a general way for a long time. Certainly all metallographers have recognized the damage that can be done to a polished surface by a piece of grit, especially if one of the softer metals, such as zinc or aluminum, is being polished. It is to be doubted, however, whether many metallographers have had any adequate idea of the particle size of the abrasives they were using or of the range of sizes encountered in any given grade of abrasive. At any rate, the descriptions of abrasives for metallographic work found in the literature are generally very meager. It was with the idea of securing an abrasive with a limited range of particle size that a study of this problem was commenced. Commercial abrasives, even when recommended for metallographic work, are made up of particles of varying sizes. One such abrasive contains some particles so fine that they remain in suspension for days and other particles so coarse that they settle at the rate of 3 in. per minute. It is evident that the coarsest particles are over 100 times the diameter of the smallest ones, or over 1,000,000 times the volume. The hopelessness of avoiding scratches under these conditions is evident. Vanderwilt1 has described a laboratory method for grading abrasives in which the elutriator is used. Le Chatelier's method, described by Goerens,2 also makes use of an elutriator. However, the elutriator is too slow for this purpose and requires too delicate a control for the range of sizes desired. The method of stirring up the abrasive in water, allowing it to settle for as given time, then drawing a measured depth off the top of the suspension, has been finally adopted as the simplest proce-dure. The largest particle in a given portion is. then determined by Stokes' law:3

Jan 1, 1931

By William H. Baer, George M. Carlton, Blake M. Loring

IT is a well known fact that age-hardening alloys remain in a supersaturated, or partially supersaturated, condition only for limited periods of time at temperatures below the solvus. In order to develop optimum properties by subjecting the quenched alloy to precipitation or age-hardening treatment at a specified temperature, it is necessary that the cooling rate be sufficiently high to maintain the required degree of supersaturation of the solid solution. In practical applications the size and shape of the specimens obviously have considerable influence on the rate of cooling and there is need for more information on the effects of various rates of cooling. Cooling rates of quenched aluminum alloys generally have been established by the type of quenching medium, the temperature of the medium and the dimensions of the specimen.1,2,3 Except for the work of Roth3 little quantitative information is available on the measurement of cooling rates of aluminum alloys and the significance of the influence of various cooling rates upon the mechanical properties. The purpose of the present investigation was to study the use of the Jominy test for aluminum alloys (Table I). EXPERIMENTAL PROCEDURE The material for this investigation consisted of extruded rods 1 ¼ to 1 ½ in. in diam. The chemical compositions were as follows: [ ] The modified L-type Jominy specimen with dimensions shown in Fig I was selected as the most convenient means for obtaining a wide range of cooling rates. The standard Jominy specimen was not satisfactory because its cooling rate was not sufficiently high. The four flat surfaces machined at right angles along the length of the L-type specimen to a depth of 0-015 in. facilitated the hardness measurements. The specimens were machined from the extruded stock with no intermediate forging. New specimens were used for each aging experiment. The specimens were given solution treatments of 2 hr duration at recommended temperatures in a Hoskins furnace (Table 2). The temperature of the furnace was maintained constant to ±3°C. No protective atmosphere was necessary because the thin film of oxide that formed on the surface of the specimens had no effect on the hardness measurements.

Jan 1, 1948

By P. W. Gilles, B. D. Pollock

THE pioneering work of Steinitz1 and Steinitz, Binder, and Moskowitz2 has shown conclusively the existence at high temperature of two additional phases in the molybdenum-boron system and thus brings to a total of six the number of structures appearing in this system. To the structures Mo2B, MOB, and Mo2B5 they have added MO3B2, a new -MOB form, and have shown that MOB,, which has the same range of composition as Mo2B5, is only a high temperature structure of the latter. This solid solution, interestingly enough, includes neither of the compositions corresponding to the stoichiomet-ric compounds, MOB, or Mo2B5, but rather at all temperatures has intermediate values of composition. These workers have also, in the course of their work, measured melting points, transition temperatures, eutectic and peritectic points in the system and have shown that Mo3B2, because of its dispro-portionation at low temperature to Mo2B and MOB, is stable only in a limited high temperature range. During the course of the present work on the vaporization properties of the molybdenum-boron compounds, a few transition temperatures were observed. When the report of the other workers appeared, it was decided to repeat, in part, their study of the system. As a result, considerable evidence has been obtained that substantiates the specific kinds of melting processes they report as well as the general features of their diagram. However, a marked difference was found between the temperatures they report and the ones observed in this study, with the latter being higher. The purposes of this paper are to present the evidence obtained in this laboratory that verifies their diagram of the system, to give some important temperatures in the system, to compare them with those previously published, and to seek an explanation of the difference. Samples The metal starting material was 400 mesh molybdenum powder with a purity stated by the manufacturer to be 99.9 pct. The initial treatment, designed to remove volatile contamination, consisted of heating in a vacuum for 10 min to a temperature of from 800" to 1000°C during which a loss of 0.3 to 0.4 pct occurred. An assay following this treatment showed it to be 99.4 pct pure, with the principal impurity probably being oxygen. The boron starting material was obtained from the Cooper Metallurgical Laboratories and the Fair-mount Chemical Co. as 325 mesh powder with manufacturers' analyses of 99 pct or better. Initial treatment consisted of heating in molybdenum in a vacuum at about 1700°C for 10 min. During this time a loss of 3.5 pct occurred. An assay following this treatment showed the different samples to have purities ranging from 95.5 to 99.0 pct with iron and carbon as the principal impurities. Following the initial treatment, the elements were combined to form stocks of Mo2B and MOB by heating pressed mixtures in a vacuum to 1100" to 1200°C to accomplish reaction and to 1500" to 1900°C for a few minutes to evaporate the more volatile impurities. Analysis of the two compounds for boron by a modification of the method of Blumenthal3 and for molybdenum by the lead-molybdate method indicated them to have purities greater than 99 pct. The individual samples to be studied had compositions in the Mo2B-MOB range and consisted of mixtures of the stock compounds. Procedure As is usually the case in high temperature work the selection of containers for the samples posed some problems. For vapor pressure studies tantalum crucibles, allowing little contact with the pressed samples, were used and some of the observations made during these experiments are pertinent to the study of the phase diagram. Most of the experiments, however, were performed in graphite containers, as were those of the previous authors. Two kinds of spectroscopic grade graphite crucibles were used. One was a % in. cylinder, 3/4 in. high, containing seven 3/16 in. holes drilled 1/2 in. deep into which were packed samples of the different mixtures weighing 250 to 500 milligrams. The other, consisting of separate crucibles, was prepared by drilling 3/16 in. holes, 1/2 in. deep into 1/4 in. graphite rods % in. long. The 7/8 in. cylinder was heated directly by induction while the small crucibles were packed in a tantalum heating element for induction heating. All heating was done in a high vacuum system in which the pressure was generally less than 1x10-5mm and never rose above 2x10-5mm when the samples were hot. The general pattern of the heating in graphite was to heat rapidly to a temperature somewhat below the desired one, then to raise the temperature slowly. The samples were held for 2 to 5 min at the maximum temperature, which in all cases was far higher than that needed to produce reaction. The short time was employed to reduce possible contamination by the crucible material and to reduce composition changes that would occur because of vaporization. After examination following the heating, the samples were reheated to a higher temperature. Temperatures were measured with a Leeds and Northrup disappearing filament optical pyrometer, certified by the National Bureau of Standards, by sighting through a window at the top of the vacuum

Jan 1, 1954