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By Thomas J. Ferree

The principle mineral concentration practices today can be divided into three basic process categories: magnetic separation, flotation and gravity separation. The basic processes have been in use for many years with only limited changes occurring in equipment design or process technique. In general, equipment and process techniques have been altered only after a need for change has developed. With exceptions for changes necessitated by new environmental laws, most equipment and process improvements have evolved in accordance with the desire to concentrate larger quantities of material at lower costs. The basic result has been the development of larger capacity equipment, to wit: larger flotation cells, larger blast furnaces, larger haulage equipment and larger chemical process equipment. However, there has been practically no changes in the size of spirals, tables, jigs, or-within limits- magnetic separators. (The exception being high-intensity magnetic separators). Consequently, a "new" device has entered the scene which may fill a void in the gravity separation field. It is called a Reichert cone concentrator.

Jan 3, 1973

By R. C. Thompson

THE lead-zinc mill of Phelps Dodge Corporation, Copper Queen Branch, Mines Division, Bisbee, Arizona, is about 3 miles from the main hoisting shafts of the Junction and Campbell mines at Lowell. All the ore treated in the mill is obtained from these two mines. The mill was designed for an all-flotation treatment of 450 tons of lead-zinc ore per day, and has been operated at that capacity since the start of milling operations on Nov. 17, 1945. An expansion program is underway whereby the mill capacity will be doubled, but this article describes only the original program and presents results of the 45o-ton operation. GENERAL DESCRIPTION The milling plant consists of an ore-receiving bin and three steel-constructed buildings housing, respectively, the primary crushing equipment, the secondary crushing equipment, and the ore-storage bins, grinding, flotation, and filtering equipment. The mill site is the one that was used for a previous copper-concentrating plant, which was operated over a period of years and then dismantled. There remained from this operation some milling equipment, a reservoir for mill-water supply, railroad tracks, tailings-disposal ponds, and a steel-constructed building, all of which were incorporated in the design of the present lead-zinc mill. The location plan of the plant layout is shown in Fig I. The flowsheet of the mill is based on the conventional scheme for lead-zinc treatment, however, several innovations were made necessary in order to utilize the existing building. Placing of equipment also was governed by the design of the building; thus making the mill unique in many respects. The following brief description of the mill building as adapted to the flowsheet points out the compactness of the installation and some of its unusual features. The first floor (Fig 2) houses the grinding equipment, with connecting conveyors for ore feeding, also an air compressor, two vacuum pumps, bins for lead and zinc concentrate storage with individual conveying systems for loading the two kinds of concentrates. Filtering of concentrates is done on the second floor, the concentrates discharging directly into the concentrate-storage bins. On this floor are the lead and zinc filters, two diaphragm pumps, sample-filtering equipment, sample-preparation room and a change room for the mill employees. The flotation machines are on the third floor, at elevations allowing a gravity flow of the lead and zinc concentrates to the thickeners. The annex to the building contains a freight elevator and provides space for all reagent mixing and feeding, also storage of reagents and grinding balls. All the water used for milling purposes is pumped from the Junction mine and

Jan 1, 1947

By R. M. Raymond

The Mining Section of The National Safety Council recently sent out a questionnaire to operators, regarding the class of rope used, specifications required or obtained, factors of safety observed, methods of fastening ropes, lubrication, inspection, and points observed when rope is to be discarded. The answers received from this questionnaire, largely from metal mines, give interesting information, from which recommendations for safety practice on hoisting ropes are drawn. Depth of Mine The depth in the majority of cases noted is from 1000 ft. to 1500 ft. (304 to 457 m.) and records come from a number hoisting from depths of 2000 to 3000 ft. Hoisting is generally done in balance, with a few companies using counterweights. Seldom, more than two shifts are occupied in hoisting. A number of large operators limit the shift for engineers to 6 hr. A few of the larger operators have separate hoists for men and supplies, with a higher class rope for men. The range of loads varies from 7 to 17 tons and greater in a few cases, with a balanced weight of 4 to 7 tons. Size or Drums and Sheaves The size of drums and size of sheaves are generally in accordance with recommendations of rope manufacturers. Only one-half of the operators use grooved drums; one, The Sulphide Corpn., New South Wales, uses wood-lagged drums. Very few operators have drums capable of holding all the rope in a single wrap. Cone drums or cone and cylinder drums, in the mines heard from, are very rare. A few mines of shallow depth use a small drum and large-sized sheave. While this may be allowed, and in some cases is all right, the size of drum and sheave should be fully equal to that recommended by the rope manufacturers and should be at least 60 to 100 times the diameter of the

Jan 1, 1923

By R. M. Raymond

The Mining Section of The National Safety Council recently sent out a questionnaire to operators, regarding the class of rope used, specifications required or obtained, factors of safety observed, methods of fastening ropes, lubrication, inspection, and points observed when rope is to be discarded. The answers received from this questionnaire, largely from metal mines, give interesting information, from which recommendations for safety practice on hoisting ropes are drawn. Depth of Mine The depth in the majority of cases noted is from 1000 ft. to 1500 ft. (304 to 457 m.) and records come from a number hoisting from depths of 2000 to 3000 ft. Hoisting is generally done in balance, with a few companies using counterweights. Seldom, more than two shifts are occupied in hoisting. A number of large operators limit the shift for engineers to 6 hr. A few of the larger operators have separate hoists for men and supplies, with a higher class rope for men. The range of loads varies from 7 to 17 tons and greater in a few cases, with a balanced weight of 4 to 7 tons. Size or Drums and Sheaves The size of drums and size of sheaves are generally in accordance with recommendations of rope manufacturers. Only one-half of the operators use grooved drums; one, The Sulphide Corpn., New South Wales, uses wood-lagged drums. Very few operators have drums capable of holding all the rope in a single wrap. Cone drums or cone and cylinder drums, in the mines heard from, are very rare. A few mines of shallow depth use a small drum and large-sized sheave. While this may be allowed, and in some cases is all right, the size of drum and sheave should be fully equal to that recommended by the rope manufacturers and should be at least 60 to 100 times the diameter of the

Jan 1, 1923

By A. Azis

The problem of expressing a nation's mineral resources in terms that convey a sense of economic reality poses a great challenge to those charged with keeping government policymakers informed. A model is presented that can be used to simulate and examine the effects of, for instance, differences in levels of output, transportation costs, mining costs or selling prices on the exploitability and expected order of development of new mineral deposits. The model in its present form can be used to screen, as a first cut, the economic from uneconomic deposits in cases where little or nothing is known about them in detail. To date, this has generally been done - reluctantly - by subjective judgments. This model provides an improvement on the "informed opinion" approach, as it forces those who use the model to be both systematic and explicit in their assumptions. Moreover, the procedure is repeat- able and each step can be defended and justified. Implications can be examined of the uncertainty associated with many of the variables. The model is entirely too general to be of use for actual feasibility studies. It is not useful, in the case of speculative or little-known deposits, to provide a model that has an input format more detailed than the information that is likely to be available. Should the user see a need for it, the model can easily be expanded to allow for the input of more detail or more variables.

Jan 1, 1977

By C. M. Nickerson

IT is apparent that repressuring of the oil measures is becoming increasingly important to the oil industry, and is a matter that warrants the best efforts of the petroleum engineer charged with applying this method of operation to a producing property. Considerable attention has. been directed to repressuring in various California oil fields, such as Dominguez and Seal Beach, which have served to sell the more progressive operator on this idea and also to keep him sold. The projects that have received most attention are in comparatively young fields where the fluid levels are high, the wells, perhaps, still flowing, where the wet gas production is of considerable volume, and where there is vital need of immediate conservation of gas, which otherwise would be blown to the air. The new State of California Gas Law, which prohibits the waste of gas, has served to interest the oil industry in -means of conserving the gas that is now being blown to the air because of lack of markets. This vitally affects the operator in the older, well-depleted oil fields of the state, because in many cases he has a limited market for gas, on account of the flush production from the more active oil fields. While the waste of gas is relatively small in these older fields, such as the Midway-Sunset area, it now attains considerable importance, since under the new State Gas Law it must be put to some beneficial use. Repressuring in the older fields of California has not received the attention of the petroleum engineers that is warranted by the present situation. This paper was prepared with the view of presenting the salient features connected with repressuring in the well-depleted oil measures of the older fields, and to show what results have already been obtained in these fields by injecting the unmarketable. gas into the sands, particularly in the Midway-Sunset field in the San Joaquin Valley of California. Other projects in the older fields of the state offer some supporting data in this connection, such as the Shields Canyon and the Brea Canyon fields.

Jan 1, 1929

By M. R. Geer Olds, H. F. Yancey

THE increase in the number of coal-cleaning plants employing dense-medium processes occurring since 1946 is especially interesting when viewed historically. Both sand and magnetite were introduced as material for heavy mediums at about the same time, sand in the Chance process in 1921 and magnetite in the Conklin process in 1922, but from that point on their records diverge. The Chance process enjoyed a steady growth from its inception, whereas no additional magnetite plants were built in the United States for over 20 years. Then, following the close of the World War 11, magnetite was again introduced, this time with marked success. During the following years some 47 plants employing magnetite medium were built. This rapid growth of dense-medium cleaning has been concurrent with widespread adoption of full-seam mining on one hand and a return to a more competitive market on the other. At a time when changing mining practice has provided cleaning plants with dirtier coal and changing market conditions have simultaneously demanded a cleaner product, the industry has through necessity turned to improved preparation. The inherently greater sharpness with which dense-medium processes can separate coal from impurity is thus helping to hold the line against ever-increasing mining costs and at the same time assisting materially in retaining badly needed markets. Although dense-medium cleaning unquestionably offers a distinct advantage when the washing problem is difficult, other methods can provide almost equally high efficiency when the coal is easy to wash. Moreover, fine coal cannot yet be treated by heavy medium in a proved process, although the Driessen cyclone is in the pilot-plant stage. Most of the present types of dense-medium equipment have been in use only a few years, and the dearth of information in the literature concerning their performance characteristics is entirely understandable, Nevertheless this information is necessary if the process is to be intelligently applied to individual cleaning problems. Without data on the efficiency of a process in a particular type of separation, it is difficult to assess the advantage to be expected from it. Similarly, the role of particle size in heavy-medium separation is important in some cases, yet there is little published information on this aspect. To mention only one more of the numerous points on which essential information is lacking, the bearing of medium characteristics on performance has been discussed only in qualitative terms. It was with the hope of providing information on some of these points that the Bureau of Mines built a dense-medium pilot plant for cleaning coal at its Northwest Experiment Station in Seattle in 1950. The plant has been operated continuously since that time, and over 50 runs have been made on 7 coals exhibiting a wide range of washability characteristics. An idea of the magnitude of this work will be gained from the fact that examination of the plant products has involved some 600 float-and-sink separations and about 2500 ash determinations. A laboratory pilot plant is especially well adapted to investigate many aspects of performance because close control over test conditions can be exercised and because a large number of tests can be made rapidly. On the other hand, factors such as consumption of medium and other cost items can be investigated satisfactorily only in a commercial plant. Actually, the two forms of investigation should be complementary, with the laboratory work pointing the way for confirming tests in commercial units. Pilot Plant The dense-medium pilot plant employed for this work comprises a 24x30-in. drum-type separating vessel, a 12-in. densifier, a 12-in. magnetic separator, a 26-in. x 9-ft vibrating screen, and the necessary pumps and conveyors for handling materials. Arrangement of these units corresponds with the flowsheet used in most commercial plants, except that a thickener is not provided for the feed to the magnetic separator. Coal and refuse discharge from the separating drum to the vibrator, which is divided longitudinally down the center. Medium draining through the first 3 ft of the screen is recirculated directly to the drum. Sprays on the middle 3 ft of the screen rinse medium from the products, and the last 3-ft section is for dewatering. Dilute medium from the rinsing and dewatering sections is pumped to the magnetic separator, where magnetic solids are recovered. These are pumped to the densifier, from which they return to the medium-drainage sump by gravity through a demagnetizing coil. The drum-type separating vessel is a scale model of a commercial unit. Feed enters axially at one end of the drum just below the surface of the bath, and float material overflows through a circular opening at the other end. Particles sinking to the bottom of the bath are picked up by lifting flights bolted to the inner wall of the drum, elevated out of the bath, and sluiced to the vibrating screen. Baffles suspended in the bath prevent float material from entering the sink-lifting flights. About 8 gpm of medium is used to sluice the feed into the drum. An additional 15 to 24 gpm, depending upon operating conditions, is added through two pipes dipping into the bath behind the baffle on the side where the sink-lifting flights enter the bath. The bath available for separation is 2 ft long and 13 in. wide, giving an area of 2.08 sq ft. Depth of bath from the surface to the top of the sink-lifting flights, measured vertically below the axis, is 6 in.

Jan 1, 1954

By G. Birkbeck, T. C. Wells

Pearlite morphology in three commercially produced, low-carbon steels has been studied using optical and electron microscopy. A reduction in the cooling rate from 600° to 6°C per hr increased the interlamellar spacing but had little effect on pearlite content. The form of the lamellae was also assessed in terms of a new parameter, the ratio of lamellae length to interhmellar spacing. The variation of this parameter with cooling rate depended on steel composition. For each steel, it would appear that an optimum cooling rate exists at which lamellar pearlite can be produced to the greatest extent, though not necessarily the same extent in different steels. PeARLITE morphology depends principally on steel composition, prior austenite grain size, and either cooling rate or isothermal transformation temperature. Alloying with manganese,'" fine grain size,374 and slow cooling rates reduce the likelihood that a well-developed lamellar structure will be produced. Instead, imperfect structures usually consisting of partly globular carbides are formed in a ferrite matrix. Such structures have been described as granular.1,2,5 degenerate,6 or semi4-pearlite. Manganese additions also progressively lower the eutectoid temperature and the carbon concentration at the eutectoid so so that, for a given carbon content. the pearlite fraction is increased as manganese is added. While the effects of manganese are understood in a general way, quantitative studies of pearlite morphology have been carried out mainly on steels of eutectoid composition;9 little quantitative data is available for low-carbon steels. Certain anomalies exist in the behavior of steels containing -1.5 pct Mn, however. The pearlite morphology in niobium-treated steel differs from that which is generally accepted as typical of low-carbon/manganese steels in that a reasonably well developed lamellar structure is sometimes observed. In the present work, the effects of changes in cooling rate from 600° to 6°C per hr on the pearlite morphology in three commercially produced, low-carbon steels have been studied by both optical and electron microscopy. I) MATERIALS The steels were of the following types: 1) 0.1 pct C, 0.4 pct Mn steel (En 2, a composition similar to specification SAE 1010); 2) niobium-treated, 0.2 pct C, 1 1/2 pct Mn steel (BS 968:1962); 3) silicon-killed 0.2 pct C, 1 1/2 pct Mn steel (En 144, a composition similar to SAE 1320). The chemical analysis for each steel is given in Table I. To facilitate the comparison of pearlite morphology in the three steels, steels 1 and 3 (which were obtained as 1/2-in. plate and 1-in.-diam bar, respectively) were hot-rolled in the laboratory to $ in. thickness, finishing at around 850°C. The BS 968:1962 was received as 1/8-in. plate finished at 885°C and was not further worked. Specimens from each steel were heat-treated in vacuo at 950°C for 1 hr and cooled at rates of 600°, 60°, and 6°C per hr. 11) EXPERIMENTAL Pearlite contents were determined by point counting (500 points) on longitudinal sections mechanically polished and etched in 2 pct nital. Pearlite morphology is usually considered in terms of interlamellar spacing. For pearlite colonies randomly distributed in space and having a true, constant interlamellar spacing So Pellissier et al.9 determined the distribution curve for the apparent spacings of the lamellae on a random section plane. The areal fraction fs of the pearlite colonies occupied by apparent spacings up to a limiting value S is given by:

Jan 1, 1969

By G. J. Ballmer

Trucks Are Valuable Auxiliary Haulage Units in Rail Pits For: ( I ) Making railroad grades (2) Establishing benches (3) Making new dropcuts

Jan 10, 1951

By John J. McCabe

INTRODUCTION Over the years, Congress has written into the Internal Revenue Code various provisions aimed at lessening at least one financial burden faced by taxpayers in the mining industry - the federal income tax burden. These provisions include the election to deduct exploration expenses, the election to defer deduction of development expenses, and the percentage depletion deduction. Careful tax planning from the very beginning of a mining venture is essential in taking maximum advantage of these tax benefits. A major decision to be made is the choice of organizational structure. (For a more detailed discussion of this subject, see McCabe, John J., "Developing Structures for Investments in Mineral Property," New York University Institute on Federal Taxation, 42 NYU 24, 1984). One such structure is the use of multiple corporations, with mining activities conducted by one corporation and nonmining activities, such as manufacturing, financing and administration, conducted by the other corporation. This technique, which will be discussed in detail below, can be particularly advantageous in delaying recapture of previously deducted exploration expenses and in maximizing depletion deductions. TAX BENEFITS IN MINING Exploration Subject to certain restrictions discussed below, exploration expenditures (i.e., expenditures paid or incurred during the taxable year for the purpose of ascertaining the existence, location, extent, or quality of any deposit of ore or other mineral, and paid or incurred before the beginning of the development state of the mine are deductible at the election of the taxpayer (Section 617(a)).* Were it not for this option, exploration expenditures would be required to be capitalized and recovered through depletion, after production begins. Exploration of a mineral deposit can be very costly and may not result in commercial production for some time. If the taxpayer has income to he offset by the deductions, a mining venture can usually be structured in such a way as to allow the offset of mining losses against the other income. With today's high interest rates, this election can be quite valuable in present value terms. Corporate and noncorporate taxpayers are subject to different restrictions on this tax benefit. Corporations may elect to deduct only 85 percent of eligible exploration expenditures (Section 291(b)). (The Tax Reform Act of 1984 reduced this to 80 percent effective for expenditures after December 31, 1984, in taxable years ending after such date). The remaining 15 percent (20 percent after December 31, 1984) is capitalized and recovered at the same rate as five-year ACRS property (Section 291(b)). Individual taxpayers are not subject to this 85 percent rule; however, any amount deducted in excess of ten-year straightline amortization is a tax preference item for individuals (Section 57(a)(5)). Individuals may avoid the tax preference by electing to capitalize exploration expenditures and amortize them over ten years (Section 58(i)). In order to prevent the taxpayer from, in effect, converting ordinary income into capital gain by selling property that has given rise to exploration expense deductions, the Code requires recapture of such deductions upon the occurrence of certain events (Section 617): 1. Upon disposition of the property, adjusted exploration expenditures are recaptured to the extent of gain realized on the disposition (Section 617(d)). 2. If the property is leased, the lessor is disallowed any depletion deduction until his adjusted exploration expenditures have been recaptured in the form of foregone depletion deductions (Section 617(c)).

Jan 1, 1985

By C. T. Bischoff

This paper describes the results of testing with colorimetric methods, using "dithizone", soil samples taken over various known copper and zinc deposits covered by glacial till. Variation in results is shown for different depths and types of overburden as well as grade and type of deposit. Limiting factors for the practical application of this work are discussed. The method has proved successful for copper and zinc deposits under clay and fine sand with depths up to at least 30 ft but fails in coarse sand and gravel unless depths are very shallow. Swamps require penetration to the underlying subsoil for sample material. DURING the past few years geochemical testing for traces of heavy metals in water, soil, rock, and vegetation has aroused increasing interest. Various techniques for field and laboratory determinations have been described in the literature, as have some of the results of application. However, little information has been published regarding application to glacial terrain. Over 90 pct of the Canadian Shield is overburdened, resulting in high costs for mineral exploration once the outcropping ore deposits have been found. This has been largely responsible for the widespread application of mining geophysics in Canada, permitting concentration of effort on area or zones considered most favorable. However, geophysical interpretation has its limitations, since at least 90 pct of the structures indicated thereby contain no valuable minerals. A more direct approach is therefore highly desirable. If underlying base metal mineralization can be detected by sampling soil or vegetation at moderate cost this should prove of great help in locating new discoveries or extending known ones. A large proportion of Canadian ore deposits contain appreciable copper or zinc, even where these are not the principal economic metals. Consequently the field of application is a broad one. This paper deals with investigations along these lines with particular emphasis on soil sampling. The problems involved from the start were: 1—Which types of glacial soil, if any, would best lend themselves to testing, and conversely which types might be considered relatively hopeless? 2—Through what depths of overburden could significant mineralization be detected? 3—What sampling techniques would prove economically applicable? This included minimum depth of sample and maximum spacing that could be allowed.

Jan 1, 1955

By A. Kenneth Schellinger

DURING the comminution of a brittle material in the presence of dry air, no known phase change or chemical reaction takes place. The energy changes associated with the comminution are those of the transformation of kinetic energy into heat energy and the energy absorption represented by the actual rupture of molecular and/or ionic bonds. The random nature of the actual grinding process in a tumbling-mill often precludes the effective application of a rupture force on each individual particle of brittle mineral present. The result must be a great deal of useless vibration on an atomic scale in both the brittle mineral and the grinding medium below the elastic limit, which vibration is eventually manifested as heat energy in the mill. The actual rupture of molecular, or ionic, bonds creates new surface on the brittle minerals; each square centimeter of this new surface then must possess a number of potential bonds, each of which could reform theoretically with the release of energy. The sum total of this potential energy of bonding per unit of new surface, minus bond satiation effects of the enclosing gas or liquid, must be the surface energy of the mineral. If the comminution of the brittle mineral were carried out within the chamber of a calorimeter, the energy released as heat could be determined rather accurately. Moreover, the total kinetic energy input to the comminuting mass (mineral and medium) conceivably could be derived by a suitable torquemeter and revolution counter if the whole calorimeter were revolved. The difference between these two energy figures should then represent the net energy (gross energy input-heat energy output) used in creating

Jan 6, 1951

By A. CHESTER BEATTY

MAGNESIUM oxide is by far the most refractory of the common oxides, since it has a melting point of 5072 deg. F. as compared with 3110 deg. F., the melting point of silica (crystobalite) ; 3722 deg. F., the melting point of alumina (corundum) ; and 4127 deg. F., the melting point of chromite (Cr2O33FeO). Its large-scale use in the lining of metallurgical furnaces dates back to the latter part of the nineteenth century, when extensive deposits of magnesium carbonate were discovered in southern Austria in the province of Styria. Before the development of Austrian magnesite, calcined dolomite, a mixture of lime and magnesia, was used in basic steel-making furnaces. The lime in the dolomite was quite objectionable, since it slaked on exposure to the atmosphere, and dolomite therefore could not be made' .into permanent furnace bottoms or into brick. The Austrian magnesite, with its very low lime content, was free from these objections and soon displaced much of the dolomite. With the extensive development of basic steel-melting, the rise of the copper industry, and other metallurgical and chemical processes which dealt with chemically basic substances, the use of magnesite grain and magnesite brick increased rapidly.

Jan 1, 1931

By Frederick Laist, H. J. Maguire

The ore from the Butte mines of the Anaconda company is quite siliceous; that is, it contains considerably less iron than is needed for the fluxing of the silica. The direct smelting of this ore, therefore, requires the addition of considerable fluxing material. The only flux available in sufficient quantity at Anaconda is limestone containing no metallic minerals. This necessarily makes the direct smelting of the ore so expensive that it is more profitable to subject the ore to a preliminary treatment by concentration. However, as the losses in the concentrator were very high, compared with the losses in the blast furnaccs, it was found, by calculation, more profitable to send ores containing more than a given percentage of copper to the blast furnaces, in spite of the higher costs of treatment. The dividing point was dependent on various factors, the chief of which was the prevailing market price of copper. It was customary, therefore, to segregate the ore as mined into two classes. Ore in the first class generally contained 5 per cent. or over of copper; this was sent directly to the blast-furnace department. Second-class ore contained less than 5 per cent. of copper, and averaged around 3 per cent.; it went to the concentrating department. With the introduction of the flotation process, it mas found possible to maintain an average recovery, as concentrates, of 96 per cent. At the same time, the cost of reverberatory smelting was very materially decreased by the use of pulverized coal. These two changes in practice made the blast furnaces unable to compete with the reverberatory furnaces on a cost basis. The distinction between first- and second-class ore was therefore abandoned, and the ore was treated by concentration. Obviously, on ores as high in iron as the Butte ores, the concentrator produces a product that is self-fluxing; in fact, it is possible to produce a concentrate containing more iron than is necessary to flux the silica. Under these conditions, the amount of basic fluxing materials required for the smelting operations became very srnall and the converter slag lost its former value as a blast-furnace flux. For this reason, a less expensive method of recovering the copper, silver, and gold contained in it than smelting in blast furnaces had to be found. It was not feasible

Jan 1, 1921

By Lillian Erskine

While brass and other copper alloys have long been listed as offering health hazards to their workers, it is questionable if the metals involved are alone responsible for the trades' records of morbidity. The entire science of foundry sanitation is the product of the past two decades; and the best standards now operative would have been regarded as little short of revolutionary prior to 1910. The special handicap to the conservation of the health of the brass founder's employees has been the fact that, unlike the iron founder, his business could be profitably carried on in almost any small and unsanitary quarters. As a result, writers on occupational disease have accepted brass chills or brass founders' ague as an inevitable feature of the trade; and respiratory disabilities, as well as phthisis and disturbances of the nervous system, are also included as prevalent among those employed. While there can be no question as to the basic accuracy of past investigations, there is every reason to doubt the necessity for an undue percentage of ill health among brass molders and casters who are safeguarded by scientific ventilating standards and by improving personal habits of sobriety. There is, however, need for a campaign of education as to the most approved practices and equipment whereby the employer may safeguard the health of his workers, and thereby reduce his labor turnover and increase his productive efficiency. Broadly speaking, copper, tin, and aluminum carry negligible health hazards. The toxic qualities of lead, however, are a menace; especially if scrap lead is recovered or bought for foundry processes. The rapid volatilization of zinc renders its fumes and flocculent condensation the chief problems in conserving the health of the brass worker. The frequent presence in the foundry air of traces of arsenic compounds, lead

Jan 1, 1919

By Hamilton Wingate

In his paper on the hydraulic mining of a low-grade gravel in California,* Mr. W. H. Radford expresses the hope that other members of the Institute will contribute, for the benefit of all, their experience in similar lines. In these notes 1 shall describe the conditions and results of the operation of the Gold Bug Mining Co., of Cleveland, O., at Georgetown, Eldorado county, Cal.

Jan 1, 1903

By V. N. Krivobok

THE study and use of low-carbon manganese steels have been curiously neglected in the general history of developments in alloy steels. Hadfield1 made an extensive study of manganese-iron-carbon alloys about 1886, which culminated in the wide application and use of Hadfield austenitic manganese steel. His alloys were nearly all high in carbon, as compared with the steels discussed in this paper, and he found a marked tendency towards brittleness in the alloys of medium and low manganese content, that is, below 7 per cent. manganese. As a result, there has been a widespread prejudice against the use of manganese steels containing between 1.0 to 3.5 per cent. manganese with comparatively low carbon. Steels of this composition have never been studied adequately in comparison with other alloy steels, and have had little use in practice. In a very recent paper by Sir Robert Hadfield2 the following statement appears: ?. . . manganese-iron alloys with little or no carbon do not seem to offer any practical employment; they only become valuable in the presence of quite high percentages of carbon, 0.90 to 1.50 per cent." The results of the present investigation are so contrary to the findings of Hadfield that the writers feel fully justified in presenting the preliminary report incorporated in this paper.

Jan 1, 1927

By Charles F. Herbert

In February of this year at the Washington, D.C., hearings of the Interior Department on the Alaska pipeline, the author was amazed to hear speaker after speaker oppose not only pipline construction but any development whatsoever in the State of Alaska. In one case, a man objected to development in Alaska because his neighbor in Maryland had cut down some shade trees. Another man recited at length from the poems of Robert W. Service. While it seemed inconceivable that such thoughtless opposition should have any effect, a check with the offices of some congressmen showed that mail was largely against any development of Alaska. In the barrage of words and redundant phrases directed against Alaska, it is difficult to find a thought worthy of study. Fortunately, there is a statutory guide towards resolution of the problem of development versus non- development. This law, the Environ- mental Protection Act of 1969, has indeed given much of industry the jitters; copper smelters see no way to comply with the new sulfur- emission standards, the Alaska Department of Highways estimates that the Act has nearly doubled the cost of planning and engineering for highway construction. In spite of these and many other troubles and costs, the Act provides a sound legal basis for defeating nihilistic emotions and, since Alaska is relatively undeveloped, adoption of environmental safeguards will be simpler than correcting longstanding abuses.

Jan 1, 1971

By Raja V. Ramani, Ralph W. Barbaro

This paper presents an application of a Mixed Integer Programming (MIP) model to the problem of locating a centralized coal preparation plant and determining the production schedule for the mines. The system used for illustrating model application is hypothetical but realistic. It consists of three mines, two potential preparation plant sites, heavy media processing plant, and one market. The goal is to maximize the total profit of the system while satisfying the coal delivery schedule over a three year period meeting both quantity and quality requirements (ash, sulfur, and heat content). In addition to the base case, four different scenarios are analyzed. These four modifications are: (1) removing the flotation circuit from the preparation plant and sending all of the fines directly to the clean coal, (2) lowering the quality of coal from one of the mines, (3) applying discounts to the unselected site to improve its competitive position, and (4) adding a new mine. A detailed discussion of the base case and the effect of these four alternatives is presented.

Jan 1, 1983

ROBERT E. TALLY,* Jerome, Ariz. (written discussion ?).-Mr. Sherman's paper, entitled "Measures for Controlling Fires at the Copper Queen Mine," embodies two new and very important features: First, automatic closing of doors communicating with the dangerous area; second, the use of a portable storage-battery fan. His other procedures, such as giving a general fire alarm, getting workmen to the surface, etc., are equally important, but have long been in use in other places. Fire and ventilating doors in connections leading to shafts, air raises, and other large timbered areas, have been in use at other mines; but the method described by Mr. Sherman for operating these doors by compressed air, through a valve on the surface, is, so far as I know, original, and should be installed in every mine where there is danger of fire. Nearly all large mines are now using artificial ventilation, with the result that there are strong air currents in the shafts and other important workings. The most important feature in mine fires is a quick and positive control of the ventilation, for it takes but a few moments for smoke to travel over a large area, with the usual results of suffocation, and destruction of property. A strong air current in an upcast shaft acts as a draft in case of fire, and unless quickly cut off the fire would soon get beyond control. An upcast shaft usually changes to a downcast when large amounts of water are used. Human lives are in more danger from downcast currents, as the smoke travels into the workings.

Jan 4, 1918