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By Enid C. Plante, K. L. Sutherland

Practical metallurgists are unanimous in stating that oxidation of mined sulphide ore adversely affects separation of the constituent minerals under standard conditions in a mill. Frequently, the need for different operating conditions in different plants is attributed either to the extent of oxidation of the mineral surfaces or to soluble oxidation products. The investigations reported in this paper study the effect of oxidation on the primary condition necessary for flotation, namely that of adhesion of an air bubble to the mineral. Oxidation Products from Sulphide Minerals Weinig and Carpenter1 state that the relative rates of oxidation of some sulphides are: FeAsS > FeS2 > CuFeS2 > ZnS > pbS > Cu2S tetrahedrite. They also show that oxidation is rapid at first, that sulphides of older geological formations are more stable than those of recent date and that mixtures oxidize more rapidly than any one mineral alone. The more rapid oxidation of mixtures is borne out by results of Gottschalk and Buehler2 who found that PbS, zns and cus oxidized 8 to 20 times faster when pyrite was present. Only the end products of oxidation, sulphate and the metal ion, were identified in these tests. Gaudin3 states that pyrite is difficult to oxidize in the laboratory. Winchell,4 using pyrite containing 99.98 pct FeS2, allowed water to percolate through a bed of this ground material. At the end of three months only a trace of Fe+++and So4- were found. Even after 10 months the quantity of Fe+++ was only 7.8 mg and SO4- 25.5 mg per liter of percolating solution. Car-michael5 confirmed Winchell's results on the slow oxidation of pyrite. Usually the oxidation of sulphide surfaces has been conducted in neutral or acid solutions, the acidity being caused by the oxidation products. Eliseev and Nagir-nyaka found that treatment of pure pyrite with lime gave only thiosulphate, but considered it possible that thiosulphate was obtained not by oxidation of the pyrite itself, but by oxidation of calcium disul-hide. Mitrofanov7 suggested that thiO~ulphate was formed by the following process: oxidation of sulphide to give free sulphur (this frequently happens with marcasite), followed by oxidation of the sulphur to sulphur dioxide and reaction of the sulphur dioxide with sulphur to give thiosulphate. Ferric and ferrous hydroxides, or similar compounds, were shown to be present. The Research Staff of the United Verde Copper Co.8 treated ore from the United verde mine with lime. Thiosulphate, sulphide and polysulphides were stated to be

Jan 1, 1949

By C. Chongo, L. Little, Q. Mclennan

Since commissioning in 2005, First Quantum Minerals’ Kansanshi Mine in Zambia has produced over three million tons of copper and 1.8 million ounces of gold from a deposit with highly varied copper mineralization. The processing plant is complex with three independent flotation circuits, gold gravity concentration, atmospheric and high pressure leaching, solvent extraction, and electro-winning. For the first ten years of operating the plant, copper mineralogy was inferred from total copper and acid soluble copper assays, and various misconceptions developed about the mineralogy which impacted operational strategy and project decisions. In 2015, an on-site mineralogy section was set up with an Automated Scanning Electron Microscope, the Zeiss MinScan. This instrument has provided detailed mineralogical data that have debunked some of the mineralogy myths that had developed, highlighting the key underlying issues affecting processing performance. This has enabled the development of projects targeting losses of specific copper minerals in defined size classes, unlocking significant value. This paper describes some of the early misconceptions around mineralogy at Kansanshi processing plant, outlining projects that have benefitted from understanding copper mineral distribution. Learnings from Kansanshi are to be applied at other FQM sites, with a similar mineralogy program being set up at the Cobre Panama project. INTRODUCTION The deposit Kansanshi Mine is located in the North Western Province of Zambia, and it has been the largest copper producer in Africa for over a decade. The deposit occurs within a broad northwest trending antiform that can be traced for 12 km, hosted in a zone of complicated faulting, abundant vein injection, breccia development and down-dropped rock units that lies within Kansanshi’s mining license. The copper mineralization within the deposit is predominantly vein-hosted, although some is strata-bound, occurring in thin bands and veinlets parallel to foliation, and some is disseminated associated with albite-carbonate alteration (Gray, 2015). Supergene enrichment has led to significant variety in the copper mineral speciation. This alteration is complex, and has not followed the text-book principles with a neat layer of supergene mineralization overlying fresh hypogene ore. Rather, the occurrence and extent of oxidation and supergene enrichment is highly irregular throughout the deposit, sometimes occurring at significant depth in localized pockets and along fault lines. Gold mineralization tends to follow the copper within the deposit, but it is dominated by coarse-grained, free-milling native gold. The extent of copper oxidation – measured as the fraction of total copper that assays as acid soluble – is used to classify ore into three main ore types, named Sulphide, Mixed, and Oxide, which are then fed to three different primary crushers for processing in independent circuits.

Jan 1, 2019

By J. Charles Miller

NATURAL carbon dioxide has recently been exploited in the United States in consequence of oil and gas developments in the Western States and the growing demand by transcontinental and transoceanic shippers for solid carbon dioxide. The discovery by drilling of commercial deposits resulted in consideration of its geographic and geologic distribution and occurrence. Its wide distribution in nature is evidenced by its occurrence in the atmosphere, hydrosphere and lithosphere. The question of origin of carbon dioxide involves consideration of theories similar to those concerning natural hydrocarbons, although available evidence favors an inorganic rather than an organic source. Areas favorable to development of commercial production appear to be those in proximity to igneous rocks or in the vicinity of decadent volcanic activity, where, it has long been known, the gas is evolved either free or dissolved in the waters of fumaroles, geysers or mud pots. Structure contour maps are presented showing the Wagon Mound anticline in New Mexico and the Farnham anticline, Utah. A well drilled on the latter supplies gas to a 20-ton plant manufacturing solid carbon dioxide at Wellington, Utah, on the Denver and Rio Grande R.R., whence it is shipped to surrounding territory and to the West Coast. Before funds are invested in a plant for the manufacture of liquid or solid carbon dioxide, whether supplied by wells or springs, an effort to ascertain the volume of gas available is advisable. The use of present methods of estimation of gas reserves with modifications to account for the quantity absorbed by any water present in the producing formation is recommended, The Farnham anticline is used as an example and cornputationa are made of the probable volume of gas originally present in the structure. No attempt was made to compute reserves in the Niland area, California, for the reason that in such areas it is believed the volume is indeterminate. INTRODUCTION Carbon dioxide, in liquid and solid form, has found a wide variety of uses other than its primary application in carbonating beverages. It is now used in more than 40 different ways, ranging from cooling or freezing to use in gas or liquid form for mechanical purposes, chemical industry, explosive, fire extinction, and tire manufacturing. Undoubtedly the chief supply of carbon dioxide is obtained from industrial plants or artificial sources and until recently natural deposits of the gas have received little consideration. Few will question the utility and perhaps dependability of artificial sources, which possess the advantage over

Jan 1, 1937

This is a sampling of the papers presented at the conference, the proceedings can be bought from the SAIMM The technique reported here involves measuring the air pressures generated within sealed boreholes located behind blasts. Results from twelve free face blasts in competent rock environments a represented. There were no instances of high pressure gas penetration, even at distances corresponding to less than one blast burden. Rapidly generated underpressures were measured in most holes, with the magnitudes of the peak pressures decreasing with distance behind the blasts. It has been proposed that the under pressures are the result of new volume created due to crack formation and overall rock mass dilation. The experimental evidence is consistent with this, including a correspondence of the onset times of the peak underpressures with the nominal firing times of the nearest blastholes. New post-blast cracks were observed when using borehole video cameras in the monitoring holes. A calculation of new volume produced within the monitoring holes is proposed as an indicator of damage. While the data exhibit considerable scatter, there is a power function dependence on scaled distance, which resembles that of typical vibration data. The data also lie within the normally quoted vibration regimes for rock damage and are consistent with measured extents of damage. It is concluded that this technique may provide an inexpensive and practical comparative measure of near-field rock damage due to blasting. It is further concluded that gas penetration is not a common phenomenon. Damage appears to be mainly attributable to the nearest blastholes and the control of charge mass in perimeter blastholes is emphasized for damage minimization.

Jan 1, 2000

By Norman Weiss

The need for more efficient utilization of labor in the metal-mining industry has been the subject of several recent editorials in the mining press, and one attractive possibility for such improvement lies in the designing of future milling plants particularly with economy of manpower in mind. We are accustomed to think of mill design as the selection and coordinated arrangement of process equipment, but if labor is to be efficiently utilized it must be given early consideration in planning the new mill. There are three good reasons for economizing in labor: (I) to reduce costs, not only in the form of wages but also in housing, hospital, welfare, and other obligations that a company accepts when a new man is added to the payroll; (2) to prepare for recurrent periods of manpower shortage, during which a well-designed plant can manage to get along with skeleton crews; (3) because a milling plant that must depend upon an inordinately large labor force for its operation and maintenance is certain to be inefficient in other respects, for many of the shortcomings in design that make necessary the excessive use of labor are generally not corrected merely by having too many men. A simple example of this is a poorly designed ore-feeding system : if this is inadequate to the extent that one man must be constantly unclogging feed chutes, it may help the situation to put two men or three at this task instead of one, but the feed rate will still be inconstant. First of the factors that affect labor saving is the location of the mill. This is often determined by transportation costs alone, and as a result the mill may be placed close to the mine regardless of the nature of the terrain. The designer must then adapt his layout to the conditions, favorable or not. This happens often at small properties, where capital investment may seem more important than operating expense. In most instances, however, a mill site carefully selected for the advantages it gives the designer in laying out the plant is the most economical in the long run, even if it involves a slight increase in cost of ore haulage. Given some freedom of choice, the design engineer may be able to find just the piece of ground that provides the proper gradient for the needs of the flowsheet, permits installation of essential yards and trackage, and perhaps supplies an excellent site for a tailing dam. Further justification for having placed the mill not too close to the mine sometimes emerges only years later as an opportunity to treat ores from other properties. When it is known from the beginning that the new mill will serve a number of scattered mines, a free choice of mill site is naturally stimulated. The most important single factor in limiting the labor requirements of a mill is its materials-handling system for bring-

Jan 1, 1947

By R. W. Gurry

In the course of a series of measurements of the rate of diffusion of carbon in austenite at about 960°C. (1760°F.) and III0°C. (2030°F.), it became necessary to determine carbon concentration when austenite is saturated with graphite, because of indications that the hitherto accepted value of the solubility of graphite in this temperature range is too low. Specimens of carbonyl iron were saturated at temperature in a hydrogen-toluene atmosphere which was in equilibrium with graphite. The carbon conteDt of the steel was measured either by combustion or by the loss of weight after substantially complete decarburization in an atmosphere of hydrogen saturated at room temperature with water vapor. The absence of carbide or graphite as a phase at temperature was demonstrated by microscopic examination of saturated specimens that had been drastically quenched, The mean result of concordant ments-—taeats-is-^that gamma iron saturated with respect to graphite contains 1.39 per cent carbon at 957°C .(i755°F.) and 1.89 per cent at III0°C. (2030°F.). These two values, plotted along with what appears from the literature to be the best value (0.69 per cent) at the iron-graphite eutec-toid temperature (738°C. or I360°F.), yield a curve that is almost linear, and which, continued upward through a short interval, leads to a limiting solubility of 1.98 per cent at the iron-graphite eutectic temperature II35°C. (2075°F.) instead of the commonly accepted value of about 1.7 per cent. Experimental Procedure The samples of carbonyl iron, with the exception of those for microscopic examination, were cylinders 3.8 cm. long, 0.475 cm. in diameter. They were placed in a graphite holder and suspended in a vertical furnace, the temperature of which was controlled precisely by means of the resistance of the platinum winding.' In the empty furnace the range of temperature over a distance of 4 cm. at the heat center was 9°C. (I6°F.) but at a given location the temperature remained within +o.5°C. (o.9°F.); it was measured by inserting a platinum-rhodium couple calibrated at the melting point of palladium, 1555°C. Since the presence of the sample probably improved the uniformity of temperature, it was decided to consider the average temperature of the sample to be 3°C. lower than the maximum measured in the empty furnace with the same setting of the controller. Data were obtained in the vicinity of two temperatures, 960°C. (I760°F.) and III0°C. (2030°F.). The samples for microscopic examination, 2.2 cm. long, 0.475 cm. in diameter, were suspended individually on pure iron wire; for these no temperature correction was applied. The carburizing gas passed into the furnace was dry, oxygen-free hydrogen saturated at room temperature with toluene; this quickly produced on the

Jan 1, 1942

By Norman Weiss

THE need for more efficient utilization of labor in the metal-mining industry has been the subject of several recent editorials in the mining press, and one attractive possibility for such improvement lies in the designing of future milling plants particularly with economy of man-power in mind. We are accustomed to think of mill design as the selection and coordinated arrangement of process equipment, but if labor is to be efficiently utilized it must be given early consideration in planning the new mill. There are three good reasons for economizing in labor: (I) to reduce costs, not only in the form of wages but also in housing, hospital, welfare, and other obligations that a company accepts when a new man is added to the payroll; (2) to prepare for recurrent periods of manpower shortage, during which a well-designed plant can manage to get along with skeleton crews; (3) because a milling plant that must depend upon an inordinately large labor force for its operation and maintenance is certain to be inefficient in other respects, for many of the shortcomings in design that make necessary the excessive use of labor are generally not corrected merely by having too many men. A simple example of this is a poorly designed ore-feeding system: if this is inadequate to the extent that one man must be constantly unclogging feed chutes, it may help the situation to put two men or three at this task instead of one, but the feed rate will still be inconstant. First of the factors that affect labor saving is the location of the mill. This is often determined by transportation costs alone, and as a result the mill may be placed close to the mine regardless of the nature of the terrain. The designer must then adapt his layout to the conditions, favorable or not. This happens often at small properties, where capital investment may seem more important than operating expense. In most instances, however, a mill site carefully selected for the advantages it gives the designer in laying out the plant is the most economical in the long run, even if it involves a slight increase in cost of ore haulage. Given some freedom of choice, the design engineer may be able to find just the piece of ground that provides the proper gradient for the needs of the flowsheet, permits installation of essential yards and trackage, and perhaps supplies an excellent site for a tailing dam. Further justification for having placed the mill not too close to the mine sometimes emerges only years later as an opportunity to treat ores from other properties. When it is known from the beginning that the new mill will serve a number of scattered mines, a free choice of mill site is naturally stimulated. The most important single factor in limiting the labor requirements of a mill is its materials-handling system for bring-

Jan 1, 1946

By James W. Halley

"Fine-grained" steels have been standard products for many years. This paper describes an investigation of the effects of some of the more common grain-growth inhibitors used to produce these steels. The properties, response to heat-treatment, and the deoxidation practice necessary to produce "fine-grained" steels have been the subjects of numerous papers. It has been shown that a substantial quantity of aluminum is necessary to produce this type of steel, and it has been indicated that titanium and zirconium produce a similar result. For some time there was a lively argument between those who felt that aluminum refined the grain and those who felt that aluminum oxide was the effective agent. This argument subsided with an apparent victory for those promoting aluminum oxide. Most of the work on grain size has divided steels into two broad classifications based upon McQuaid and Ehn's original carburizing test at 1700°F or some similar test at this temperature. If the steel coarsened at this temperature, it was coarse grained, while if it did not it was he grained. Quantitative data concerning how much a certain amount of aluminum raised the coarsening temperature and affected the properties, all other factors being constant, are very meager. Similar information regarding titanium and zirconium is practically nonexistent. The published information on aluminum yields some quantitative information. Epstein, Nead and Washburn' gave the first quantitative information on the amount of aluminum necessary to prevent grain growth at 1700°F. They showed that the addition of a pound to a pound and a half of aluminum per ton (0.05 to 0.075 per cent) would assure steel that would not coarsen during a McQuaid-Ehn test at 1700°F. Herty, McBride and Houghg reported that a steel containing 0.020 per cent acid-soluble aluminum coarsened between 1742° and 1832°F., while steels containing 0.016 per cent and 0.017 per cent acid-soluble aluminum coarsened between 1652° and 1742°F. compared with similar aluminum-free steels that started to coarsen at 1580°F. Herty3 also showed a marked improvebent in low-temperature notched impact resistance with increasing acid-soluble aluminum up to 0.030 per cent. McQuaid4 compared the coarsening of steels containing 0.011 to 0.204 per cent total aluminum and found a maximum coarsening temperature at 0.098 per cent aluminum. He had no samples between 0.015 and 0.098 per cent aluminum. Houdrement6 found a marked decrease in coarsening temperature with aluminum contents above 0.50 per cent. Kinzel, Crafts and Egan6 found that aluminum and zirconium improved the low-temperature impact resistance of a low-alloy steel, while titanium was detrimental. Sims7 shows that sulphur increases the coarsening temperature and that the maximum coarsening temperature is between 0.04 and 0.06 per cent added aluminum. The fact that the coarsening temperature passes through a maximum with increasing aluminum content appears to be well established. The aluminum content that

Jan 1, 1947

By Walter S. Lutz, J. T. Griffith

The rainfall during the last six months of 1942 in Scranton and vicinity was 24.06 in.—only 4.35 in more than the average for this district in any equivalent period—yet that rain forced The Hudson Coal Co. to handle 38.8 tons of water for every ton of coal mined. It filled to overflowing, or to a higher level, 14 flooded mine areas holding, in total volume, 8.3 billion gallons of water in storage; brought about a pumpage of 11.4 billion gallons in the months of December 1942 and January, February and March 1943 Half of the total energy generated or purchased (44,322,980 kw-hr.) was required to keep the mines sufficiently free of water to permit the mining of coal during a period of much needed output. Fig. I shows the extent to which this flooding of exhausted and idle properties has proceeded, and is known to be proceeding further, at the upper end of the Northern field. This rainfall (24.06 in.) means 80,559,444 tons of water on approximately 46.2 sq. miles of area tributary to the pumping plants. During the high rainfall period, 72,687,661 tons went underground, being pumped at the stations in this territory. This means an inflow of 88per cent, a runoff of 12 per cent compared with 64 per cent during normal conditions in a normal region. During this period no floods, either surface or underground, had to be dealt with, although the streams and rivers were high on several occasions. In general, the rainfall for the first three months of 1943 can be described as heavy but not abnormal.9 These figures not only indicate the large volumes of water to be handled but, by their very size, they also emphasize that the mine-water problem is not to be conquered, at least from an economical standpoint, by merely buying and installing more and more pumps. Into the solving of this water problem must go much planning, adhering to the principle of keeping the highest percentage of water at the highest possible level for the longest period of time; the storage of huge volumes in sumps capable of being reduced after the flood inflow and at rates less than the inflow; plans for prevention of surface inflow; provision of the minimum of pumping capacity arranged to as nearly 100 per cent pumping load factor as is possible; intercompany joint pumping plants; the conducting of water from one plant to another and from one property to another over distances measured in miles; priorities in power to the threatening situations; and. balance of the necessities and dangers existent in one colliery over another, day by day, when quantities of water are running high. Despite these astronomical figures of water volumes, their disposal to the surface is accomplished reasonably cheaply because, although the annual cost of $575,020 is a large sum, the cost per ton is about 0.1168 cents. Credit for this goes

Jan 1, 1947

By Norman Weiss

The need for more efficient utilization of labor in the metal-mining industry has been the subject of several recent editorials in the mining press, and one attractive possibility for such improvement lies in the designing of future milling plants particularly with economy of manpower in mind. We are accustomed to think of mill design as the selection and coordinated arrangement of process equipment, but if labor is to be efficiently utilized it must be given early consideration in planning the new mill. There are three good reasons for economizing in labor: (I) to reduce costs, not only in the form of wages but also in housing, hospital, welfare, and other obligations that a company accepts when a new man is added to the payroll; (2) to prepare for recurrent periods of manpower shortage, during which a well-designed plant can manage to get along with skeleton crews; (3) because a milling plant that must depend upon an inordinately large labor force for its operation and maintenance is certain to be inefficient in other respects, for many of the shortcomings in design that make necessary the excessive use of labor are generally not corrected merely by having too many men. A simple example of this is a poorly designed ore-feeding system : if this is inadequate to the extent that one man must be constantly unclogging feed chutes, it may help the situation to put two men or three at this task instead of one, but the feed rate will still be inconstant. First of the factors that affect labor saving is the location of the mill. This is often determined by transportation costs alone, and as a result the mill may be placed close to the mine regardless of the nature of the terrain. The designer must then adapt his layout to the conditions, favorable or not. This happens often at small properties, where capital investment may seem more important than operating expense. In most instances, however, a mill site carefully selected for the advantages it gives the designer in laying out the plant is the most economical in the long run, even if it involves a slight increase in cost of ore haulage. Given some freedom of choice, the design engineer may be able to find just the piece of ground that provides the proper gradient for the needs of the flowsheet, permits installation of essential yards and trackage, and perhaps supplies an excellent site for a tailing dam. Further justification for having placed the mill not too close to the mine sometimes emerges only years later as an opportunity to treat ores from other properties. When it is known from the beginning that the new mill will serve a number of scattered mines, a free choice of mill site is naturally stimulated. The most important single factor in limiting the labor requirements of a mill is its materials-handling system for bring-

Jan 1, 1947

By F. M. Stewart, F. H. Callaway, R. E. Gladfelter

In this paper oil initially in place is calculated by the various methods commonly used for analysis of water drive fields using data available as time progresses. Rate and pressure are predicted by means of the same methods using data available at the end of two years' history. These predictions are then compared with subsequent performance permitting an evaluation of the methods, which are Modified Schilthuis, Simplified Hurst, van Everdingen et al (1952), and the Electric Analyzer. INTRODUCTION Several variations of the material balance equation, differing in the term used to evaluate water influx, are in common use for estimating oil in place, predicting rate and volume of water influx, and forecasting producing rates and reservoir pressure in water drive reservoirs. The electric analyzer with an electrical network set up to simulate the oil reservoir and aquifer and their contained fluids may be used for the same purpose. Examples of the application of one or another of these methods have appeared in the literature. The accuracy and reliability of the material balance for calculating oil in place have been covered.' However, there have been few articles in which more than one method has been applied to an actual field allowing a comparison. It is felt that a field example, simple in nature, where the Modified Schilthuis, Simplified Hurst, van Everdingen et al (1952), and the Electric Analyzer have been applied would be of benefit in providing an evaluation of these methods as to their comparative usefulness and reliability in calculating initial oil in place and predicting performance. The Torchlight Tensleep reservoir is a small field in a large aquifer. The reservoir oil is greatly under-saturated, having a bubble point of only 55 psi and 2 cu ft/bbl of gas in solution. Changes in the reservoir volume factor in the range of pressures encountered are small, meaning little error in converting production to reservoir conditions. The pressure has declined from 1,618 psi initially to 137 psi during the first four and one-half years. Water production has been negligible. For Torchlight, the material balance equation as discussed later is considerably simplified by the omission of terms having to do with free gas. Thus with simpler equations, less variables to consider, the large pressure drop, the absence of a gas phase, and rapid pressure-rate adjustments, which mean an unusual amount of history in a short time, the reservoir appears suitable for comparing the methods listed above. GENERAL INFORMATION ON THE FIELD The Torchlight field is located on the eastern side of the Big Horn Basin, Big Horn County, Wyo., approximately 3 miles east of the town of Basin. The Tensleep reservoir was discovered in September, 1947. It lies at a depth of about 3,000 ft and produces oil of 35" API gravity. A total of five wells have been completed in the Tensleep, but only four are now producing. The reservoir is well defined in areal extent in that a total of 13 wells have penetrated the Ten-

Jan 1, 1955

By W. Jacobi, F. Steinhausler, A. C. James

INTRODUCTION In 1979 the Committee on Radiological Protection and Public Health of the Nuclear Energy Agency (OECD) defined a programme of work needed to improve radon dosimetry and monitoring. A Group of Experts in this field was established by the NEA and at the first meeting of the Group the authors were designated an initial task with the following terms of reference:- To review dosimetry of radon, thoron and their daughters, and in particular:- - to examine analytical models for the assessment of dose to individuals who are exposed to aerosol mixtures of these nuclides; - to examine the state of knowledge of the characteristics of aerosol mixtures of these nuclides and other factors that influence the dose to the lung, and identify those parameters which should be measured for dosimetric purposes in radiological surveillance programmes; - to identify areas where further study or research are needed. In this presentation we attempt to examine, with these objectives, the relationship between exposure to radon and thoron daughter aerosols expressed in terms of the working level month l (WLM), or its SI equivalent, and dose accumulated by sensitive tissues of the lower respiratory tract. Over the past two decades many attempts have been made to calculate the lung dose equivalent to l WLM exposure in mine atmospheres, usually in terms of the maximum dose to basal cells in a small part of the bronchial tree. The calculations have employed variously detailed models of airway anatomy, aerosol deposition, ciliary clearance and dosimetry of tissue at the bronchial surface, and are well reviewed elsewhere (Walsh, 1970; UNSCEAR, 1977; Jacobi, 1977; Fry, 1977). They agree in the conclusion that basal cells in segmental bronchi (between about the third and fifth bronchial generations) receive the highest dose, although estimates of this dose differ by more than an order of magnitude, ranging from about 0.2 to 12 rads per WLM (2 mGy to 120 mGy). The most critical parameters of these models are the assumed depths of basal cells below the bronchial surface and the fraction of the WL carried by free ions. When more realistic values of these parameters are substituted (Jacobi, 1977), the dispersion in calculated maximum doses to basal cells is reduced to give a much smaller range from about 0.3 to 1 rad/WLM (3 to 10 mGy/WLM). The parameters and methods employed in these various models, however, remain too diverse to infer the sensitivity of the dose conversion to practical variations in aerosol characteristics and the additional factors determining lung deposition and clearance. We have therefore taken the approach of developing newer models (Jacobi and Eisfeld, 1980; James et al, 1980) that can encompass the range of available experimental data on the physical and biological behaviour of radon and thoron daughters and also accommodate the different interpretations that can be put upon them. We have paid particular attention to the averaging of dose over tissues of equal sensitivity in the lung

Jan 1, 1981

By R. W. Gurry

In the course of a series of measurements of the rate of diffusion of carbon in austenite at about 960°C. (1760°F.) and III0°C. (2030°F.), it became necessary to determine carbon concentration when austenite is saturated with graphite, because of indications that the hitherto accepted value of the solubility of graphite in this temperature range is too low. Specimens of carbonyl iron were saturated at temperature in a hydrogen-toluene atmosphere which was in equilibrium with graphite. The carbon conteDt of the steel was measured either by combustion or by the loss of weight after substantially complete decarburization in an atmosphere of hydrogen saturated at room temperature with water vapor. The absence of carbide or graphite as a phase at temperature was demonstrated by microscopic examination of saturated specimens that had been drastically quenched, The mean result of concordant ments-—taeats-is-^that gamma iron saturated with respect to graphite contains 1.39 per cent carbon at 957°C .(i755°F.) and 1.89 per cent at III0°C. (2030°F.). These two values, plotted along with what appears from the literature to be the best value (0.69 per cent) at the iron-graphite eutec-toid temperature (738°C. or I360°F.), yield a curve that is almost linear, and which, continued upward through a short interval, leads to a limiting solubility of 1.98 per cent at the iron-graphite eutectic temperature II35°C. (2075°F.) instead of the commonly accepted value of about 1.7 per cent. Experimental Procedure The samples of carbonyl iron, with the exception of those for microscopic examination, were cylinders 3.8 cm. long, 0.475 cm. in diameter. They were placed in a graphite holder and suspended in a vertical furnace, the temperature of which was controlled precisely by means of the resistance of the platinum winding.' In the empty furnace the range of temperature over a distance of 4 cm. at the heat center was 9°C. (I6°F.) but at a given location the temperature remained within +o.5°C. (o.9°F.); it was measured by inserting a platinum-rhodium couple calibrated at the melting point of palladium, 1555°C. Since the presence of the sample probably improved the uniformity of temperature, it was decided to consider the average temperature of the sample to be 3°C. lower than the maximum measured in the empty furnace with the same setting of the controller. Data were obtained in the vicinity of two temperatures, 960°C. (I760°F.) and III0°C. (2030°F.). The samples for microscopic examination, 2.2 cm. long, 0.475 cm. in diameter, were suspended individually on pure iron wire; for these no temperature correction was applied. The carburizing gas passed into the furnace was dry, oxygen-free hydrogen saturated at room temperature with toluene; this quickly produced on the

Jan 1, 1942

By James W. Halley

' FINE-GRAINED " steels have been standard products for many years. This paper describes an investigation of the effects of some of the more common grain-growth; inhibitors used to produce these steels. The properties, response to heat-treatment, and the deoxidation practice necessary to produce "fine-grained" steels have been the subjects of numerous papers. It has been shown that a substantial quantity of' aluminum is necessary to produce this type of steel, and it has been indicated that titanium and zirconium produce a similar result. For some time there was a lively argument between those who felt that aluminum refined the grain and those who felt that aluminum oxide was the effective agent. This argument subsided with an apparent victory for those promoting aluminum oxide. Most of the work on grain size has divided steels into two broad classifications based upon McQuaid and Ehn's original carburizing test at 1700°F. or some similar test at this temperature. If the steel coarsened at this temperature, it was coarse grained, while if it did not it was fine grained. Quantitative data concerning how much a certain amount of aluminum raised the coarsening temperature and affected the properties, all other factors being constant, are very meager. Similar information regarding titanium and zirconium is practically nonexistent. The published information on aluminum yields some quantitative information. Epstein, Nead and Washburn' gave the first quantitative information on the amount of aluminum necessary to prevent grain growth at 1700°F. They showed that the addition of a pound to a pound and a half of aluminum per ton (0.05 to 0.075 per cent) would assure steel that would not .coarsen during a McQuaid-Ehn test at 1700°F. Herty, McBride and Hough2 reported that a steel containing 0.020 per cent acid-soluble aluminum coarsened between 1742° and 1832°F., while steels containing 0.016 per cent and 0.017 per cent acid-soluble aluminum coarsened between 1652~ and 1742°F. compared with similar aluminum-free steels that started to coarsen at 1580°F. Herty3 also showed a marked improvement in low-temperature notched impact resistance with increasing acid-soluble aluminum up to 0.030 per cent. McQuaid4 compared the coarsening of steels containing 0.011 to 0.204 per cent total aluminum and found a maximum coarsening temperature at 0.098 per cent aluminum. He had no samples between 0.015 and 0.098 per cent aluminum. Houdrement5 found a marked decrease in coarsening temperature with aluminum contents above 0.50 per cent. Kinzel, Crafts and Egan6 found that aluminum and zirconium improved thy low-temperature impact resistance of a low-alloy steel, while titanium was detrimental. Sims7 shows that sulphur increases the coarsening temperature and that the maximum coarsening temperature is between 0.04and 0.06 Per cent added aluminum. The

Jan 1, 1946

By V. V. Valleroy, A. J. Cornelius, B. T. Willman, G. W. Runberg, L. W. Powers

This paper reports the results of an investigation into the use of steam as a recovery agent. High oil recoveries by steam, as much as 100 per cent greater than by water flood, were demonstrated in laboratory linear cores. The principal mechanisms responsible for the additional oil have been identified as (1) thermal expansion of the oil, (2) viscosity reduction and (3) steam distillation. Although the first two also enhance recovery by hot water flood, they are effective with steam injection at lower volumes of produced fluids. Steam-zone residual oil saturation is independent of initial saturation. Recovery by steam is greater for lighter oils because they contain a greater fraction of steam-distillable components. Equations for the steam-zone advance in a radial homogeneous system are presented; these show the effect of reservoir parameters. The displacement in the hot liquid region moving ahead of the steam front can be estimated by a calculative procedure which is suggested. The method which takes account of the temperature gradient is applicable also to a regullar hot water flood. INTRODUCTION Considerable industry interest in the application of heat to petroleum reservoirs as a means of increasing oil recovery has been evident in the past few years. Three thermal methods which have attracted interest are (1) the injection of saturated steam, (2) the injection of hot water and (3) the injection of air into a formation to support combustion within the reservoir. Much has been published in recent years on underground burning, and there have been several papers on hot water injection. Although there are several early reports1,2 on steam injection, no recent experimental studies have been published on this recovery method. In our early thermal methods research, a marked improvement was found in oil recovery by steam injection compared with recovery by water injection in the same system. An investigation subsequently was undertaken to obtain quantitative information from which to evaluate steam injection as an oil recovery method. In addition, it was desired to investigate the recovery mechanism of hot water as compared with that of "cold" water and steam injection. An understanding of the mechanisms by which these processes operate would generally define favorable field conditions for their successful application. This paper reports the results of the laboratory study and suggests calculative approaches to estimate the field behavior of the steam-injection process. EXPERIMENTAL METHOD EQUIPMENT A schematic diagram of the equipment used for the displacement tests is shown in Fig. 1. Essentially, it consisted of a core with suitable means for recording temperatures, a feed system for injecting steam or cold or hot water, and a receiving train for condensing, cooling and measuring produced fluids. For the steam experiments, air-free saturated steam was piped from the boiler to the core face in insulated lines that were provided with steam traps and bleeds to remove any condensate formed in the lines due to heat losses. A 12-in. section of the line immediately ahead of

By J. T. Griffith, Walter S. Lutz

The rainfall during the last six months of 1942 in Scranton and vicinity was 24.06 in.—only 4.35 in more than the average for this district in any equivalent period—yet that rain forced The Hudson Coal Co. to handle 38.8 tons of water for every ton of coal mined. It filled to overflowing, or to a higher level, 14 flooded mine areas holding, in total volume, 8.3 billion gallons of water in storage; brought about a pumpage of 11.4 billion gallons in the months of December 1942 and January, February and March 1943 Half of the total energy generated or purchased (44,322,980 kw-hr.) was required to keep the mines sufficiently free of water to permit the mining of coal during a period of much needed output. Fig. I shows the extent to which this flooding of exhausted and idle properties has proceeded, and is known to be proceeding further, at the upper end of the Northern field. This rainfall (24.06 in.) means 80,559,444 tons of water on approximately 46.2 sq. miles of area tributary to the pumping plants. During the high rainfall period, 72,687,661 tons went underground, being pumped at the stations in this territory. This means an inflow of 88per cent, a runoff of 12 per cent compared with 64 per cent during normal conditions in a normal region. During this period no floods, either surface or underground, had to be dealt with, although the streams and rivers were high on several occasions. In general, the rainfall for the first three months of 1943 can be described as heavy but not abnormal.9 These figures not only indicate the large volumes of water to be handled but, by their very size, they also emphasize that the mine-water problem is not to be conquered, at least from an economical standpoint, by merely buying and installing more and more pumps. Into the solving of this water problem must go much planning, adhering to the principle of keeping the highest percentage of water at the highest possible level for the longest period of time; the storage of huge volumes in sumps capable of being reduced after the flood inflow and at rates less than the inflow; plans for prevention of surface inflow; provision of the minimum of pumping capacity arranged to as nearly 100 per cent pumping load factor as is possible; intercompany joint pumping plants; the conducting of water from one plant to another and from one property to another over distances measured in miles; priorities in power to the threatening situations; and. balance of the necessities and dangers existent in one colliery over another, day by day, when quantities of water are running high. Despite these astronomical figures of water volumes, their disposal to the surface is accomplished reasonably cheaply because, although the annual cost of $575,020 is a large sum, the cost per ton is about 0.1168 cents. Credit for this goes

Jan 1, 1947

By J. G. Mravec

The SO-called gas-tube method as developed by Hare, Peterson and Soler for determining the type and content of gases in molten steel is particularly adapted for determining the hydrogen content in molten steel. The development of this method has been quite completely discussed in a paper published by these co-workers in 1937.' The purpose of this paper, therefore, is to familiarize with this method those who arc not well acquainted with it. The principle of the gas-tube method is as fol1ows: One end of an evacuated gas tube is sealed with a very thin copper disk. when this end of the tube is immersed in a test spoonful of molten steel, the copper disk melts almost immediately, allowing a portion of the steel to be drawn up into the evacuated space. Any dissolved gases, and those produced by reactions between the components of the steel during freezing, are liberated into the unfilled portion of the sampling tube. A blob of steel that forms on the immersed end of the tube after the sample is drawn up seals up the end, making it airtight again. The evolved gases are subsequently drawn out of the gas tube through a suitable valve and analyzed by means of an Orsat apparatus. The volume of the sampling tube is sufficiently large so that a pressure considerably below atmospheric is produced after the gas evolution. The low pressure is necessary to effect a nearly complete removal of gases from the steel. Construction oF Sampling Tube The sampling gas-tube apparatus is illustrated in Fig. I. Seamless S.A.E. tubing used for section A is first cut into 7-ft. lengths. Through cxperiencc it was found necessary to use two sizes of tubing for this part, "is-in. and 3/4-in., respectively. The smaller tubing is generally used on the hotter or lower carbon steels. If the inside wall of the tube A is coated with grease or oil, it should be burned out before being machined as follows: One end is machined on the inside diameter to a depth of 34 in., so as to form a seat for a 7/8-in copper disk, which is stamped out from 0.005 to 0.007-in. thick copper sheet. The other end is machined on the outside diameter to a diameter of 1 5/16 in., to take a small reservoir or trap member I?. About 4 in. below the same end, a 1/2-in. hole is drilled to receive the 1/2-in. copper tubing leading to the large reservoir C. The inside of tube A is then thoroughly cleaned with a round wire brush, such as is generally used to clean out shotgun barrels. It is very important that the surface be absolutely free of rust or any nonmetallic coating. Next, the thin copper disk is soldered in place with low-melting silver solder.

Jan 1, 1945

By G. E. McElroy

THOUGH air jets occupy only a minor role in the ventilation of mines and other enclosures, their importance in this respect is greater than is commonly recognized and may be considerably increased as our conception of their modes of operation becomes clearer. The object of this paper is to discuss some of the ways in which air jets are involved in mine ventilation and some possible applications to the ventilation of mines and other enclosed spaces. The mention of air jets ordinarily suggests the use of some form of an air injector, such as the Venturi blower occasionally used for auxiliary ventilation in mines, or the Saccardo (injector) system used to a limited extent for ventilating railroad tunnels. Actually, the term covers such common phenomena as the discharge of air from fan pipes and coal-mine line brattices, and even from smoke stacks and fans. A knowledge of the general mode of air entrainment by jets also helps to explain certain phenomena-some mine-gas explosions, for instance-that otherwise would remain obscure. Jet action has been involved in three recent series of Bureau of Mines experiments in which the author has participated: experiments on fan-pipe discharge, on the dilution of gases discharged from a stack and on the ventilation of a ship's hold. Available data on the discharge of jets into large openings and on the design and performance of air injectors have been analyzed in considerable detail and supplementary experiments are planned for after the war. The results of the fan-pipe experiments have been combined with the analyses of data on jet discharge into large spaces, and have been published.1 The results of the stack-gas dilution tests, which refer to a jet discharged into a crosswind, have also been published2 and a paper on the design and performance of injectors has been prepared for publication.3 These technical experiments and analyses form the background and the documented evidence for the following nontechnical discussion of the use of air jets in mine ventilation. JET CHARACTERISTICS The jets commonly encountered are those in which the air is in turbulent-as opposed to laminar, or streamline-motion. The characteristics under discussion apply only to such turbulent jets and have their basis in theories of the mixing of turbulent air currents. Such jets have basic characteristics that are present, though variable in form, regardless of whether the jet expands freely in space, whether the expansion is limited by adjacent walls, or whether the jet propels air through a flow circuit. They expand at definite angles and entrain the surrounding air, even though the entrained air must be obtained by recirculation of their own air streams. As the quantity of air in motion in the mixed stream increases, the average velocity of the stream decreases. In the case of entrainment in the free

Jan 1, 1945

By Howard R. Keil

PRIOR to World War II, approximately 500 tons per day, or one-third of the ore being shipped to the Golden Cycle Mill at Colorado Springs, Colo., from the Cripple Creek district, was being treated in the flotation plant. The remaining 1000 tons, along with the flotation concentrates, received the usual roasting and cyanide treatment. Since it was not possible to make a disposable tailing on the general run of Cripple Creek ores by flotation alone, it was necessary to treat the flotation tailings by cyanidation, along with the roasted ore. By this treatment, an overall economical tailing could be made. The gold ores from the Cripple Creek district are the highly siliceous sulphotellurides with approximately 3 pct total sulphides. In the lower grade ores most of the values are found in very finely disseminated pyrite. In some of the higher grade ores the tellurides, calaverite and sylvanite, can be recognized. After several years of floating Cripple Creek gold ores, having a very low mineral content, it was concluded that this type of ore required a Type of flotation machine that would give a rather violent aeration and agitation. This action produced a fairly deep froth column which contained not only the free mineral, principally pyrite, but also most of the middling product; this was most essential in producing a low grade tailing by flotation alone that could be discarded to waste. After investigating most of the current types of flotation machines, it was decided to convert to the Fagergren type. During World War II, the flotation plant was modified to handle lead and zinc ores. The treatment of these ores was carried on until April 1946. At this time the flotation plant was again reconverted to handle the Cripple Creek ores. SELECTING ORES FOR FLOTATION During the period of reconversion a considerable amount of experimental work was carried on in the laboratory to determine if a disposable tailing could be made by flotation alone, on certain selected ores from the Cripple Creek district. It was found that most of the dump ores and a few of the low grade mine ores could be treated by this method with very satisfactory results. After the preliminary test work in the laboratory, it was decided to carry on further work on a plant-size scale to see if similar results could be obtained. On April 15, 1946, a part of the plant was put into operation as a pilot plant. The various ores that had been found suitable for treatment by this method in the laboratory, were run through the pilot plant at the rate of 200 tons per day for a period of several months. Results were so encouraging that it was decided to put the plant in full-scale operation. Since that time plant capacity has been increased from 500 to 1000 tons per day. This has been done not by the installation of more machinery but by (I) crush-

Jan 1, 1947

By J. B. Hess, R. L. Dietrich

In the mechanical twinning of magnesium on the {1012} planes the crystal-lographic deformation is such that, in the direction of the hexagonal axis [0001], twinning is possible only undef tension stress, while in the directions perpendicular to the hexagonal axis, twinning is possible only under applied compression. Thus in extruded magnesium and magnesium alloys where the extruded textures are best described as either [1010] or [1120] fibers,' the basal planes are all nearly parallel to the extrusion axis and such material is susceptible to ( 1or2 ) twinning by compression in the direction of the extrusion axis. In order to explain the results of cyclic stressing in alternate tension and compression upon the tangent elastic modulus in such material, Dorn and Thomsen2 have postulated that, following twinning under applied compression, a retwinning or un-twinning occurs as a result of residual microtensile stresses immediately upon removal of the applied compressive force. Recently, Carapella and Shaw4 have used this supposition of untwinning in magnesium in discussing their results on indentation hardness and on cold drawing of MI sheet. In trying to confirm this hypothesis by suitable direct experiments, the authors have been unsuccessful in detecting any evidence whatsoever that untwinning follows the release of applied compression and are forced to conclude that this hypothesis is improbable or, at least, that the quantity of material involved is small. For the experiments, small cylindrical specimens 0.040-in. diam. X 0.120- to 0.140-in. long were etched from 0.0625-in. extruded rod. The ends of the specimens were accurately faced off parallel in a jeweler's lathe. The compression jig was simply an ordinary set of micrometer calipers with an anvil interposed between the spindle and the specimen. The anvil was free to move axially but restrained from rotating so that the rotation of the spindle would not introduce a twisting stress on the specimen. With this arrangement the compressive strains could be estimated to 0.0001 in. and X ray pinhole photographs could be taken with or without load as required. Fig I is the pattern of an uncompressed specimen of AZ3I alloy, showing the approximate [1010] extrusion fiber. The X ray beam is normal to the extrusion axis; in this position it is particularly sensitive in detecting twinning if one directs his attention to the diffraction ring from the basal plane. Since twinning on (1072 planes produced an 86" reorientation of the crystal lattice, the original basal reflections occur at the 3 and 9 o'clock positions while the (0002) reflections from any twinned material will occur at the 6 and 12 o'clock positions. Fig 2 shows the same specimen compressed 3.0 pct, photographed while still under load. This strain has been just enough to produce a detectible amount of twinning and the sensitivity is roughly indicated by the fact that no alterations were produced in the intensities or maxima of the other diffraction rings. A strain of

Jan 1, 1949