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By Frank A. Hill

A visitor to the Northern anthracite coal-field in 1844 wrote: " History and song have hallowed the valley of Wyoming, and everything pertaining to it seems wrapped in an atmosphere of romance." Song and history have since been almost forgotten in the rush of enterprise and business energy, while the invested capital which has built 160 colliery plants, producing last year 17,031,826 tons of coal and giving employment to 50,000 men and boys, lends an atmosphere the inspiration of which is the very sustenance of prosperity and progress. The fundamental cause of this great valley's financial strength, however, rests on the fulness of its mineral wealth. Its mountains surround and its cities are built upon anthracite coal. Its foundation, physically, financially and socially, is coal. Its geology has governed its growth and developed its greatness. In 1768, ten years, prior to the celebrated Wyoming massacre, two Connecticut Yankees, named Gore, settled near Wilkesbarre, Luzerne County, and, burning Pennsylvania anthracite in their blacksmith's forges, were the first to utilize it as a fuel. Its first use in the household was also at Wilkesbarre, where, in 1808, Judge Fell burned it in an open grate. Carbondale, Wilkesbarre, and Scranton were early recognized as the coal centers of the field.

Jan 1, 1887

By David A. Keys

The object of this investigation was to determine with a vertical magnetic variometer the extent of the titaniferous ore deposit that occurs not far from Ivry in Terrebonne County, Quebec, about 70 miles north of Montreal. The Ivry mine, on lots 37 and 38 of range V, Beres-ford township, Quebec, is one of the largest and most accessible to transportation of the deposits of ilmenite in Quebec. The ore consists of an aggregate of ilmenite anhedra serving as host for numerous fine plaques of hematite. No magnetite is known from any occurrence near this lot, and in this way the deposit differs from many others that contain a large percentage of magnetite with the ilmenite. The deposit at Ivry thus has much lower magnetic susceptibility than those containing abundant magnetite, and a sensitive magnetometer was necessary for the work, although a body of magnetite with ilmenite about a mile from the Ivry mine was successfully prospected with a dip needle. The principal orebody is exposed in a quarry known as the Ivry mine, from which about 16,000 tons of ore have been shipped. Small exposures of ilmenite isolated by drift serve to show the presence at the surface of some of the bodies indicated by the magnetometer, but most of the hill is covered with glacial till and second growth hardwood. At the lower levels the till gives place to fluvioglacial sands and gravels from which the forest has been cleared. The increasing thickness of the sand and gravel as the hill descends makes the tracing outward of the deposits from the rocky knob of anorthosite a difficult matter. Within the orebodies too, some masses of anorthosite persist, which may account for some local irregularities. Fig. 1 is a photomicrograph, taken by reflected light, of a typical specimen of the ore. The dark patches are hematite and the light parts are ilmenite No trace of magnetite was found. Pig. 2 is a photomicrograph of a thin section of the country rock showing the texture of the anorthosite. The mean of four analyses quoted by Robinson1 is: titaninm, 19.26 per cent; iron 45.41 per cent. The ilmenite ore is not sufficiently magnetic to deflect an ordinary compass needle, but the effects of the induced magnetism in the ore by

Jan 1, 1940

By David A. Keys

The object of this investigation was to determine with a vertical magnetic variometer the extent of the titaniferous ore deposit that occurs not far from Ivry in Terrebonne County, Quebec, about 70 miles north of Montreal. The Ivry mine, on lots 37 and 38 of range V, Beres-ford township, Quebec, is one of the largest and most accessible to transportation of the deposits of ilmenite in Quebec. The ore consists of an aggregate of ilmenite anhedra serving as host for numerous fine plaques of hematite. No magnetite is known from any occurrence near this lot, and in this way the deposit differs from many others that contain a large percentage of magnetite with the ilmenite. The deposit at Ivry thus has much lower magnetic susceptibility than those containing abundant magnetite, and a sensitive magnetometer was necessary for the work, although a body of magnetite with ilmenite about a mile from the Ivry mine was successfully prospected with a dip needle. The principal orebody is exposed in a quarry known as the Ivry mine, from which about 16,000 tons of ore have been shipped. Small exposures of ilmenite isolated by drift serve to show the presence at the surface of some of the bodies indicated by the magnetometer, but most of the hill is covered with glacial till and second growth hardwood. At the lower levels the till gives place to fluvioglacial sands and gravels from which the forest has been cleared. The increasing thickness of the sand and gravel as the hill descends makes the tracing outward of the deposits from the rocky knob of anorthosite a difficult matter. Within the orebodies too, some masses of anorthosite persist, which may account for some local irregularities. Fig. 1 is a photomicrograph, taken by reflected light, of a typical specimen of the ore. The dark patches are hematite and the light parts are ilmenite No trace of magnetite was found. Pig. 2 is a photomicrograph of a thin section of the country rock showing the texture of the anorthosite. The mean of four analyses quoted by Robinson1 is: titaninm, 19.26 per cent; iron 45.41 per cent. The ilmenite ore is not sufficiently magnetic to deflect an ordinary compass needle, but the effects of the induced magnetism in the ore by

Jan 1, 1940

By Frederick M. Becket, Russell Franks

WideSpread experimentation has been conducted in recent years to devise a means of preventing intergranular corrosion in austenitic chromium-nickel steels of the 18 per cent chromium-8 per cent nickel type. Additions of titanium to these steels have been used commercially to overcome this dificulty, as advocated by Krupp of Germany, and the latest literature reference concerning improved details of this method is the article by Bain, Aborn and Rutherford.' Extensive work by the present authors has shown columbium also to be a particularly valuable agent for inhibiting intergranular corrosion. This paper does not deal with steels that are largely or wholly austenitic; it is limited to a discussion of the results obtained through the use of titanium and columbium in plain high-chromium steels. For many years titanium has been used commercially by some steelmakers as a cleanser, and to a lesser extent as a hardener, but then in greater percentages. The results of the work to be described disclose that titanium and columbium in proper Proportions act as effective softeners in the high-chromium steels. It is well known that chromium intensifies the hardening effect of carbon in steel; it imparts pronounced air-hardening properties within a rather wide range of chromium content; it provides increased ductility in steels of given strength and hardness; and it enhances resistance to corrosion and oxidation at elevated temperatures to a marked degree. However, in applications requiring a combination of some of these valuable properties great hardness is frequently a detriment rather than a merit. When the objective is a soft and ductile steel having optimurn corrosion resistance, the producer strives for a carbon content as low as commercially practicable. Despite these efforts some of the commercial high-chromium steels low in carbon present considerable difficulties during production, owing to their air-hardening properties, and in consequence the economic result is not favorable. A good example is found in the 5 to 6 per cent chromium steels employed chiefly in the oil-cracking industry. Sleels

Jan 1, 1934

By H. M. Otte, A. G. Crocker

SLIP, twinning, stacking faults, and precipitates on well-defined planes in a crystal produce traces that are visible on either a polished or an etched surface. The purpose of this note is to establish the usefulness and limitations of the information obtained from such traces when determining the orientation of single crystals or grains in a polycrystal. Attention will be confined specifically to traces produced by octahedral planes in cubic crystals, though most of the remarks can be taken as more generally valid. The problem of determining the crystal orientation is completely solved in principle' and in particular for the special case under consideration.1'2 Tables are available3 in which the orientations can simply be looked up when given the two or three angles between the traces. Thus a quick and simple method is available for obtaining the orientation of a crystal from surface traces. In this note two examples will be given from recently published papers and will show the use of the method with the aid of the Tables, as well as some of the precautions that should be taken in the procedure. As part of an analysis of X-ray Kossel lines, Peters and Ogilvie 4 determined from them the crystallographic orientation of the specimen. They chose a Cu-5.10 wt pct Si alloy that had been partly transformed from the a (fee) phase to the k (hcp) phase. The latter formed as very straight platelets along the octahedral planes of the a phase, and a photograph (Fig. 7 of Ref. 4) was reproduced in the paper4 of the region for which the orientation had been determined from the Kossel lines. Apart from a very small area (about 0.15 mm diam) near the top center of the photograph, the region shown was clearly all of one orientation. Traces in four different directions were readily identified. Starting with the smallest angle and proceeding counterclockwise, the angles between the traces were measured to be 27° 43', 55° 16', 64 3 43 1, and32"26'. These add up to 1802 08 1, indicating a cumulative error of 8'. The correct values should, of course, be taken from the original glass plates, or at least from prints on distortion-free paper, so the precision of the quoted values is actually greater than their accuracy. The solutions can be looked up in the Tables3 in four ways, depending on how the angles are paired, tion for the minutes of arc is required and the solution obtained will be identical for all the four pairs if the values of the angles are not in error. In a practical situation, such as this one, there are several sources of error, so that the four solutions obtained will all be slightly different. An arithmetic average of the four solutions is mathematically incorrect, but may, in special cases, give an answer close to the correct one. A rigorous procedure for refining the orientation has been described by Mackenzie;5 he gives a least-squares method that can be applied, if necessary, to this type of problem. The orientation thus determined can be compared with the one from the Kosel lines and discrepancies larger than the experimental error assigned to nonor thogonality of the X-ray or electron beam to the specimen. Here the example will only be used to illustrate the general procedure for making a very rapid rough check. From measurements on the stereographic projection, Fig. 9 of the paper,4 we can determine that the surface normal is approximately (0.434, 0.704, 0.562). From the Tables3 we find, using 28°, 55" as a good approximation to the first pair (27°43', 55°16'), that the surface normal is (0.453, 0.678, 0.579).* These two values are within 2 deg of each

Jan 1, 1967

By P. D. I. Honeyman, P. A. Faust

At the Miami plant of the International Smelting Co., Inspiration, Ariz., there are four reverberatory furnaces, all 120 ft. long. At present only one of these furnaces is in operation. It is operated as a wet-charge furnace, the charge being distributed for about 65 ft. down the furnace length. Usually this charge contains about 12 per cent moisture and runs 25 per cent sulfur, of which there is eliminated in the furnace from 25 to 30 per cent. The furnace is oil-fired, the fuel being burned in a battery of Cananea-type, low-pressure burners. The flame is kept short and conditions so adjusted as to obtain complete combustion within the furnace. The average furnace draft, taken through the verb arch, amounts to only 0.02 in. of H2O. Draft is purposely kept low to cut down infiltration through the flues and boiler settings, and control dampers are regulated to maintain this uniform condition throughout the charging as well as the smelting period. Hot gases discharge from the furnace uptake through a common header flue to the waste-heat boilers. The general layout and dimensions of the plant are shown in Fig. 1. Four Stirling boilers, two of class M-26 (713 hp.) and two class M-31 (850 hp.), serve the furnace. Three usually are sufficient to take care of normal operation, the fourth boiler being used as a spare when necessary. The distribution of hot gases to the various boilers is regulated by the dampers in the boiler exit flues. Normally the boilers closest to the furnace uptake would evaporate more water than do those farther away, but dampers are regulated so as to distribute fairly well the load between the several boilers. Steam from the waste-heat plant is delivered to the adjoining power house of the Inspiration company, where it is used both in the Nordberg blowing engines and for electrical generation. Steam pressures and temperatures recorded are those in the main power-plant header and are thus inclusive of all losses in about 500 ft. of 10-in. line. Steam for smelter heating and auxiliary use is taken from the main waste-heat header, is reduced to 125 lb. and distributed through separate lines. Boiler feed water is supplied by the power plant and consists of returned condensate, augmented to a large extent by make-up water

Jan 1, 1934

By G. V. Cullen, N. G. Siviour

It has previously been reported that metal loadings of up to 6.8 m. eq/g. of dry coal can be obtained by contacting brown coal or lignite with ammoniacal solutions of ammine forming metals. Application of equilibrium expressions to further experimental results indicates that the extraction is by ion exchange, that the exchange is divalent-univalent for Zn, Cd and Ni, but is mainly univalent-univalent for Cu, and that to obtain reasonably uniform values for the equilibrium constants correction formulae must be applied to compensate for multiple functionality of the coal and its pH dependence in ion exchange. The results are shown to be consistent with the hypothesis that metal ammine complex cations are exchanged for protons derived from carboxyl and/or phenolic hydroxyl groups in the coal.

Jan 1, 1981

By Wilbur C. Helt

The United States is self-sufficient in bituminous coal and lignite resources to meet the high range of energy demands forecast for the remainder of this century and beyond. The known or proved recoverable reserves of energy from coal in the United States are estimated to be 88% of all fuel reserves, whereas the energy reserves for oil are estimated to be 276, gas 376, shale oil 5% and uranium 2% (as now used). The demand for all forms of energy used in the United States is increasing at a rate of 3% to 4% per year. Presently, coal is competitive on a supply and cost basis with other fuels in conversion to electrical energy and as a source of energy for converting raw minerals to metals and nonmetals, but it is not competitive in transportation and other uses that demand the high flexibility that oil and gas enjoy because of their form. Coal conversion to fluid forms is essential to make it interchangeable with gas and oil to meet the markets they now occupy. Essentially, the chemical constituents that make up coal are the same hydrocarbons that make up oil and gas; it is a matter of reconstituting the proportions of hydrocarbons to make coal interchangeable with the other fuels. The production of coal has fluctuated widely between the peak production of 631 million tons in 1947 to a low of 403 million tons in 1961. Throughout this period, the total value of the production, f.o.b. mines, has averaged more than $2 billion annually. Beginning with 1962, there has been a steady upward trend in coal production with only minor fluctuations that were mainly caused by labor disputes. Output for the latest typical year, 1970, was 588 million tons. Production in 1971 was affected by a major strike and dropped to 548 million tons. Exports of coal from United States to foreign countries have increased steadily and developed into a substantial annual market for over 71 million tons per year. About 56 million tons of this is suitable for metallurgical purposes, and the remaining 15 million tons is exported for electric power generation. The largest single-industry demand for coal is in the electric generating industry, Fig.1. Roughly, the production of electricity has doubled every

Jan 1, 1973

By John C. Kleiber, George M. Meisel

From time to time since the initial installation of an experimental flotation section in 1940, Permanente Cement Co. has made efforts to beneficiate the low-grade limestone fraction which occurs in its deposits at Permanente, Calif. The encouraging results outlined in this article are the result of Permanente's latest effort. THE PERMANENTE LIMESTONE AND THE REASON FOR ITS BENEFICIATION The Permanente deposit is a sheet of Calera limestone over 700 ft thick lying within the Franciscan sediments of the San Francisco peninsula. Limestone suitable for cement manufacture underlies, overlies, and is interbedded with low-grade limestone and chert. Andesite intrusions are found as a secondary gangue. The limestone is identified as follows:

Jan 3, 1964

BARBOUR, PERCY E., Deputy Supt., (Captain), New York State Troopers; Captain, 22nd Regiment, N. Y. N. G. BARLING, 1i. B., 1st Lieutenant, Engineer Officers' Reserve Corps. CHAPMAN, R. H., Major, Engineer Officers' Reserve Corps. DUNN, THEODORE S., 1st Lieutenant, Engineer Officers' Reserve Corps. DWIGHT; ARTHUR S., Major, Engineer Officers' Reserve Corps. FORBES, CARROLL R., Captain, Engineer Officers' Reserve Corps. GOODSPEED,. G. E., JR., 1st Lieutenant, Engineer Officers' Reserve Corps.

Jan 7, 1917

Flaubert, as we know, laid stress on the selection of the right word, le mot juste, the precise epithet, the word that belongs to the thing. A sentence, or even a paragraph, may be spoiled. by the use of a word that is not proper, that does not fit or is foreign to the meaning intended. An Australian mining expert is reported to have said: You have a property of considerable value, and what I saw warrants s good development of very rich cubicular galena, with gossan intermixed. 'Cubicular' means 'belonging to a small bedroom', a 'cubicle' being a sleeping compartment. He meant 'cubic', of course. This is a malapropism, and reminds me of the lady from Chicago that, on her return from Europe, said how much she had enjoyed the rural parts of France, because, among other things, it was "so delightful to hear the French pheasants singing the mayonnaise to the tune of the Los Angelus". It is impossible to amalgamate coal-tar thoroughly with the pulp in the agitating-tank. In metallurgy, 'amalgamate' refers to the combination of mercury with one or both of the precious metals, forming an amalgam, which is an alloy of mercury with another metal. It differs from other alloys, such as brass or bronze, in being made without the aid of heat, inasmuch as mercury is molten at the ordinary temperature. The word comes to us from the Greek through Old French and means 'something soft'. It is a mistake to use technical terms, having a specific meaning, for purposes that are foreign to that meaning. The writer of the quotation is not discussing any combination of metals; he is dealing with the physical process of mixing and is using 'amal-

Jan 1, 1931

By R. M. Hudson, P. E. Perry

Weight-gain data obtained by nitriding low-carbon sheet steel in an amrnonia CNH,) atmosphere indicated that the process obeyed a parabolic rate law. The calculated actization energy for nitriding in the range 964" to 1268°F agreed reasonably well with published data. At 1358"F, rate data indicated that the activation energy decreased. Weight-gain data obtained by uszng mixtures of NH3 -Nz at 1268°F containzng jrom 10 to 100 zol pct NH3 also obeyed a parabolic rate law. The rate of 'nitriding increased with an increase in the NH3 content of the gas Mixture. It is well-known that steel heated in gas mixtures containing ammonia (NH3) takes up much larger quantities of nitrogen than steel heated in nitrogen, both gases having a total pressure of 1 atm;' this phenomenon can presumably be attributed to the catalytic decomposition of NH3 on the steel surface to furnish nascent (monatomic) nitrogen. This process was studied bv Brunauer. Jefferson, Emmett, and Hend-ricks at furnace temperatures of 752" and 831°F2 using mixtures of NH3 in Hz. Englehardt and wagner3 reported that, at a furnace temperature of 914°F and under their experimental conditions, both nitriding and denitriding were controlled by the rate of gas-metal reactions at a steel surface rather than by the rate of diffusion of nitrogen in iron. The present study was undertaken to obtain information on the kinetics of nitriding low-carbon steel strip at higher temperatures so that practical rates for short-time strip-annealing treatments could be estimated. Variables studied included time: temperature, and NH, content in the annealing atmosphere. Mechanical and chemical characteristics of steel nitrided in this manner will not be considered in the present article. MATERIALS AND EXPERIMENTAL WORK The samples used were from a commercial low-carbon steel, 0.0244 cm thick, in the cold-reduced condition. The chemical composition of this steel is given in Table I. Panels were cut to 5.1 by 17.8 cm, degreased in toluene, and weighed just before treatment. Four specimens were nitrided under each of the experimental conditions. A study was made of the nitriding rate of steel in a 100 vol pct ammonia atmosphere, 740 mm pressure, at five specific temperatures within the range 964" to 1358°F. The nitriding rates of steel in ammonia-nitrogen gas mixtures containing 10, 18, 26, 50, and 100 vol pct ammonia, 740 mm total pressure, at 1268°F were also determined. All atmospheres used were dried by successively passing them through drying towers packed with soda lime and with Linde Molecular sieve Type 4A. Quoted gas compositions refer to those entering the furnace. Specimens were held in the constant-temperature zone of a vertical annealing tube furnace for times of 14, 3, 5, 10, or 15 min. Gas flow rates were maintained at 3.8 cu ft per hr, which was nineteen volume changes per hour for the system used. The rate of flow was selected to provide a high level of free NH3 for cracking on the steel surface where the ammonia gas is most effectively used as a nitriding agent. The vertical annealing tube furnace consisted of a Hevi-Duty tube furnace with a 2 1/2-in.-ID mullite ceramic high-temperature tube. The constant-temperature zone (controlled within 10°F) was about 10 in. long. After each specimen was degreased, a hole was punched in one end, for attaching the specimen by hook to a chain so that it could be lowered into or raised from the high-temperature portion of the tube by means of a power-driven winch. A stainless-steel access port with O-ring seals was connected by suitable glass-to-metal seals to the cool upper portion of the furnace tube. After the weighed specimen was placed in the access port, the furnace tube was evacuated to approximately 10"3 torr, and then the system was flushed thoroughly with the atmosphere under study. When the gas flow rate and constant-temperature zone of the furnace were established, the specimen was lowered into the constant-temperature zone. The atmosphere flowed from the top to the bottom of the vertical furnace tube and was then vented. For all these runs, during the first 3 min of the time the specimen was in the constant-temperature zone of the furnace the specimen was heating up to the tempera-

Jan 1, 1970

By Bunting S. Crocker

PEBBLE grinding at Lake Shore is not a temporary wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct - 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 1 1/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to, $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change, back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1 1/4-in. steel ball charge. To change pebble size it is necessary only to change two sets-of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean. out the customary supply of grinding balls kept on hand. Also it has not always been ' possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant should be designed to use as many units-or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, .2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant. and an extremely efficient-one. In this plant the ratio of ball mills to tube mills was 1: 2. When the much cheaper pebble mills were substituted for the tube mills, this. ratio -was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was. proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on -.8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.1 This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without .grate discharges-both with 1 1/4-in. and 3/4-in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report' was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3 1/2-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are, lengthy and may be covered in a separate 'report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The -8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands .from the bowl sent to the primary pebble mills, see Fig: 1. In the pebble mill circuit the ore is ground to 90 pct -325 mesh (24 pct + 28 microns) In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO2, 80 pct insoluble. Its grindability at different. meshes has been shown near the top of 'the list in F. C. Bond's grindability tests' Lake Shore is not shown, but an adjacent-mine with identical ore, Wright-Hargreaves, is. Reasons for the-Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1952

By J. P. Witherow

The plant of Messrs. Oliver Brothers and Phillips, the only one in operation until January, 1886, has not been available for any further experiments since those of Mr. R. W. Hunt, described in his paper on the subject, at the Chattanooga meeting. Continuous working has, however, led to important improvements in plant, and furnished valuable results in practice. The difference between the converter and plant of 1884, as illustrated in our Transactions, vol. xiii., p. 759 and following pages, and those of 1885, may be observed in Plates 1. and 11. accompanying this paper. Plate I. gives the elevation and plan, in half-section, of one of our latest 3-ton converters. Plate 11. gives a plan of the general arrangement of a plant adapted to turn out slabs, blooms, and billets for the general market. The letters on Plate I. refer to the following parts :

Jan 1, 1886

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By H. W. Camp

The past year has had no revolutionary change in the process of petroleum refining, although there have been improvements and developments in practically every phase of operation, due largely, perhaps, to economic pressure. A few years ago, a Mid-Continent skimming plant using 35 gravity crude obtained 28 per cent. gasoline and naphtha; 9 per cent. kerosene, and the remaining yield gave gas oil and fuel oil. Figuring manufacturing cost, but not including sales, transportation, overhead and depreciation, the realization was $1.45 per bbl. of crude. Today, a thoroughly modern and up-to-date skimming plant, having all latest equipment, shows a realization of $1.74 per bbl. on the same basis. Thus it is evident that progressive developments have taken place in the last 2 or 3 years when potential earnings have been increased about 20 per cent. The average refiner is compelled to exert his utmost ingenuity in order to operate with a profitable margin. He has two factors to consider—recovery of maximum yields of profitable oils and the reduction of operating expenses. For the past 8 years, the margin between price of gasoline and cost of crude has grown narrower, constantly. Both these prices are determined by external conditions and not by the refiner himself, but the refiner might have been able to increase that margin had he more carefully controlled his output, limiting his production of light oils to current consumption. Be that as it may, this same economic pressure has given impetus to progress in improvements of primary distillation equipment, including stills, towers and heat-saving devices. A pipe still and bubble cap fractionating equipment can now replace a battery of shell stills, giving better yields, better control, elimination of rerunning and lowering of operating costs. The most recent pipe stills are being equipped either with radiant-heat furnaces or air-preheating systems and improved heat exchange equipment. These can be operated on a fuel consumption as low as 1 per cent. of the charge as compared to a fuel consumption of

Jan 1, 1929

By Jeremias K. Chitunda

Significant cost savings and improved ground stability are two initial indications from the current wet process shotcrete ground support program at the Lakeshore mine. The area of shotcrete for ground stability is not strange to the Lakeshore operation. Since the beginning of the project, various shotcrete techniques have constantly been experimented with, incorporating all the latest technological developments. Over the years, the use of shotcrete has been employed as a primary, temporary, or permanent support, singly or in combination with rockbolts and steel sets. The extent of shotcrete usage compared with steel sets and rockbolts has varied from time to time as dictated by the overall state of the art. In 1972, for example, the ratio of the linear drift footage supported with shotcrete to that supported with steel sets averaged 1.0. In 1973, that ratio was about 3.0. Overall, for the Lakeshore mine, there are about 8000 drift ft of shotcrete, 8000 drift ft of steel sets and timber, 750 drift ft of formed concrete, 21,000 drift ft of rockbolts, and 3700 drift ft of combined type support.

Jan 12, 1974

By R. W. Raymond

The death of Louis Janin, which took place Mar. 6, 1914, at Santa Barbara, terminated the earthly career of almost the last survivor, and one of the most illustrious members, of that group of American mining engineers and metallurgists who, in the last quarter of the nineteenth century, directed the development of the mineral resources of the Pacific slope. Louis Janin was born in 1836 at New Orleans. His grandfather, an officer in the French army, had emigrated to this country, with his family, at or before the outbreak of the French revolution. His father became an

Jan 1, 1915

By U. N. Bhrany, N. Arbiter

F. C. Bond (Consulting Engineer, Processing Machin-Dept., AllisChalmers Manufacturing Co., Milwaukee) This is a very comprehensive paper. It deals with 1) size distribution functions, 2) energy-particle size relationships, and 3) criteria for grinding efficiency. The method of analysis used consists of tracing the increase and decrease of successive Tyler scale screen fractions as grinding progresses. The treatment is based upon the assumption that the work done in grinding is concentrated entirely upon the coarse size fractions in the mill. It is postulated that when a feed consisting entirely of particles of one size fraction is batch ground that finer fractions increase at a constant rate as grinding progresses, until finally each in turn begins to be ground and to decrease in amount. It is shown that under these conditions the empirical Gates-Gaudin-Schuhmann log-log size distribution plot can be logically explained. It is difficult to imagine any condition of ball mill operation, or any probability analysis, which would result in no grinding contacts on the finer particles present. However, the authors' postulate must be accepted if their conclusions are to be sustained. They conclude that the equation expressing the for mation rate of each size fraction can be resolved into the G.-G.-S. size distribution equation, and that the curvilinear portion of the log-logsize distribution line represents the only particle sizes which are being ground as well as produced. The author's statement that the G.-G.-S. Eq. 4 applies only to the finer portion of the size distribution while the Rosin-Rammler Eq. 6 applies over the whole size distribution range certainly supports the view that an exponential equation of the Rosin-Rammler type corresponds more closely to actual size distribution than does the power law G.-G.-S. equation. However, the G.-G.-S. equation is used to derive Eq. 8 and Eq. 9, which state that the energy required to grind a ton through a particular mesh varies inversely as the particle size to the exponent -a , whic h is the slope of the extended G.-G.-S straight line. This is stated to apply to both batch and continuous grinding. The G.-G.-S. slope exponent thus becomes very important in all crushing and grinding analyses by this method. The authors include some data which can be used to test their conclusions, as described below: We will consider grinding a feed composed of 14 mesh (1190u) to -200 mesh (74u). The average of 17 Allis-Chalmers grindability tests on quartz gives a work index (Wi) of 12.77. The -200 mesh product should have 80 pct passing 41.6u or P = 41.6. The equivalent 80 pct passing size of the 14 mesh feed particles is considered to be the particle diameter, as is true for spherical particles. Thus F is 1190u. In the work index equation 10 wi 10 wi 127.7 127.7 P f 41.6 1190 = 16.1 kw-hrperton. In Table I the ball mill exponent for quartz is 1.0 and from Eqs. 9 and 10: w=kw-hr/ton =Constan/(do/di) = 16.1 ton so that (1190) Constant/ Constant and the Constant equals 259. The tons per horsepower-hour value is 0.0464, in agreement with Table IV. The exponents for other minerals are listed in Tabl e 11. In Table IIA, the kilowatt hours per ton required to grind these minerals between the same sizes as the quartz, both by the authors' method and the work index equation, are listed. According to the authors kw-hrperton = 259/(16.1)".

Jan 1, 1961