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By M. H. Sommer

MANY attempts have been made to develop a relation between the tensile properties and the deep-stamping qualities of metals com-monly used in deep drawing. These operations are generally performed cold and the metal is always stressed beyond the elastic limit. However, the problem cannot be readily solved without a proper investigation of (1) the relation between actual stress and actual strain in the plastic range and (2) the stresses that are set up in the metal during the process of deep stamping. Much work has been done in recent years to add to our knowledge on plastic deformation of metals and the shortcomings of the ordinary tensile test, particularly in the plastic range, have been sufficiently recognized. It has been widely acknowledged that the local contraction, which appears in a tensile test after the maximum load is reached, is not a general characteristic of the material but is peculiar to this form of testing. The plastic deformation of materials is explained by the slip of their crystals along certain planes and this rearrangement should be similar whether observed in a tensile, compression or torsion test, and if the results of any of these forms of testing show deviations, it must be due to other influences, as for example the local contraction in a tensile test. For this reason the tensile-test curve as obtained from the measurements taken on a tensile-test machine or registered by an autographic device has been supplemented by the so-called "actual stress-strain curve," which is constructed from the observed tensile-test data by eliminating the effect of local contraction and is in complete alignment with the stress-strain values obtained in a compression or torsion test. The actual stress-strain curve shows the increase of actual unit stress in relation to unit elongation beyond the yield point and

Jan 1, 1934

By S. C. Hollister

An evaluation of the function of the engineer, so that means whereby education can best serve his needs can be adopted. THERE has been a steady increase in specialized branches of engineering during recent years. Some of these are simply specializations in the function of engineering, such as welding engineering, tool and die engineering, petroleum engineering, and refining engineering. Others are associated more closely with the basic sciences, as for example, geological engineering or engineering physics. Still others tend toward the border line of business, such as administrative engineering or general engineering.

Jan 1, 1950

By M. T. Abasov

Among a host of challenges dictated by the explosively developing scientific and technical revolution, the improvement of control and management is one of the most essential problems. This problem is also of far-reaching importance for the oil and gas industries. An oil and gas extracting complex is an involved multilink system characterized by a great number of elements and by intricate interconnections between these elements. Of paramount importance for these systems is the quality of control. Any deviation of the control from the optimum reduces the effectiveness of the system and deteriorates its basic technical and economic factors. In case of large systems, even a relatively small deviation from the optimum may result in considerable material and power losses. Optimization of control of large production systems is therefore one of the basic means to enhance their effectiveness. The present paper discusses a development control problem for a multilayer gas deposit using the method of simultaneous exploitation of several strata by means of a single well. The optimum control is achieved by determining the dynamics of changes in the number of producing wells in a reservoir with regard to the depth of occurrence of these reservoirs as well as the difference in the time of introduction of the wells from drilling and their transfer to another reservoir. A mathematical model was constructed for the problem under consideration and a method was given to solve this problem. The problem was dealt with under the assumption that either each of the reservoirs or a part of them have vertically overlying areas and that there is no hydrodynamic interference between the horizons. It was

Jan 1, 1977

By H. Hubert

THE first attempts at direct utilization of blast-furnace gas in engines were made in 1895. For a considerable time the gas had been burnt in Cowper stoves for heating the blast for the furnace, and under the boilers which supplied steam to the blowing-engines, and others serving the furnaces. It was natural, therefore, that the idea of directly employing it in gas-engines should have occurred simultaneously to several engineers, notably to Lürmann and to Lencauchez, who had pointed out the blast-furnace as a powerful -gas-producer. Nevertheless, nowhere had any attempt been made to apply it to this purpose up to the end of 1894, when Thwaite proposed it to Mr. James Riley, of the Glasgow Iron & Steel Company. About the same time investigations were being made in Belgium and in Germany, independently of Thwaite's experiments, which were not generally known on the Continent. The industrial world, which up to that time had hardly favored the idea, had thus been gradually prepared to receive it. The gas-engine, long restricted to small sizes and dependent upon the use of an expensive fuel obtainable only in large centers, now began to make headway. At the Paris Exhibition of 1889 two engines of 100 h.p. were shown and excited much interest among engineers. One had four cylinders, and was made at the celebrated works of the Deutz Company, and the other was a single-cylinder engine, exhibited by two French designers, Messrs. Delamare-Deboutteville and Malandin. In the meanwhile the design of gas-producers had made important progress, completely freeing the new engine from its

Nov 1, 1906

By R. E. Allen

THERE was once a time when a practical oil man would appraise or buy a producing property on the basis of from $200 to $500 per barrel of average daily settled production. Curtailment-has, for the present at least, caused a virtual abandonment of the "barrel -of settled production" as a yardstick of, oil property value and in its place has come the "barrel of allowable production," which, although an artificial thing, is nevertheless now the principal factor governing the rate of ?pay out? or investment return. The potential production of a well, settled or otherwise, is of less importance in determining the present worth of a well than is the rate at which it may be produced. Speaking only of true potential for the moment, it is obvious that of two similar wells having the same potential, the one having the highest rate of allowable production would be worth proportionately more than the one of lesser allowable. The discount factor for deferred return and risk can easily vary the present values of the two wells by as much as 20 per cent when allowable rates are not greatly different.

Jan 1, 1934

By A. J. Mcintosh

PRICE changes in competing fuels in the last three decades have reflected the changes in the consuming habits of the people of the United States. Prior to World War 11 the importance of fuel oil and fuel gas increased almost constantly while the relative importance of both bituminous and anthracite coal decreased. During this same prewar period coal prices increased to a much greater extent than oil prices. In fact, in many years, oil prices declined and in most cases are now no higher than they were back in 1925-26. During the war natural economic trends were entirely upset because of the tremendous petroleum requirements of the Army and Navy. Under these conditions many industrial plants were converted to the use of coal and the new installations in home heating and industrial plants of all kinds which might have used oil were practically eliminated. Although petroleum prices were rigidly controlled by OPA coal prices increased several times during the war. An over-all statement as to the rela-

Jan 1, 1947

By F. J. Chernosky

Most previous studies of coal flotation utilized chemically pure reagents. Since such reagents are not available in quantity, a study of various wmmercially available reagents as frothers was undertaken. Flotation procedures similar to those used in a large number of commercial coal flotation plants were employed. The reagents used were compared on yield-selectivity-wst basis. The selectivity was found to be relatively constant with yields above 70% but the reagent cost varied greatly. The types of reagents available were compared. In the past, a number of papers have been written on the use of various reagents as coal flotation frothers. For the most part, these have been an attempt to correlate the intrinsic chemical and physical proper- ties of the reagents or reagent groups to the floata-bility of various fine coals. Because of the type or nature of evaluations made, chemically pure reagents were normally used. Generally, these reagents are unavailable in commercial quantities. Therefore, the results obtained with these reagents were usually limited in usefulness to the plant operator or metallurgist. This paper presents the results of a study in which commercially available forms of various reagents were used as flotation frothers. Flotation procedures which would conform in general to those employed in a large number of commercial coal flotation plants were utilized. The objective of the test was to compare these reagents on yield-selectivity-cost basis, that is, 1) the yields obtained, 2) the ash levels obtained and 3) the reagent cost relating to these results. REAGENTS The reagents employed have been grouped into the following chemical classes: 1) aliphatic alcohols, 2) terpines, 3) glycols (synthetic water-soluble

Jan 1, 1963

By JAMES GAYLE

(Presented at the Washington meeting, May 3, 1905, and simultaneously sent to the Iron and Steel Institute, for presentation at the meeting of that Society in London, May 11, 1905.) IT is to be regretted that the data respecting the use of dry-air blast, which was presented to the Iron and Steel Institute at its meeting in the United States, in October, 1904, should. have been restricted to the period from August 25th to September 9th inclusive, and from September 17th to 30th inclusive. In order to present the paper at all it was found necessary to limit the record of operations to the period above stated. In the discussion of the paper it would appear from the conclusions of some participating therein, that they have been placed at some disadvantage in considering the economies obtained through the use of dry-air blast, by reason of the data covering such a short period. It is the purpose of this communication to present in detail the record of operations of the Isabella Furnaces from November, 1904, to March, 1905, inclusive, as shown by the furnace-records. The month of October, 1904, is not included, as the furnace using dry-air blast was stopped several times for repairs, and while the record was a very good one considering these interruptions, yet it was not a continuous one. The record for Furnace No. 1 using dry-air blast, and the comparison with Furnace No. 3 using natural air for the month of November, 1904, is set forth in Table I. By a reference to my former paper it will be noted that the month of November is the beginning in this country of the winter period, when the atmosphere decreases rapidly in humidity. The month of October being the transition month in the autumn and April in the spring, and furnaces using natural air approach more closely, from November on, to the conditions obtained by the use of dry-air blast. The data given in

Jul 1, 1905

By Stanley W. Ivosevic

The South-Central Osa Peninsula produced around 30,000 kg (100,000 troy oz) of gold since prehistoric times. Gold freed by erosion from quartz lodes in the Mesozoic Nicoya ophiolite complex was reconcentrated through three generations of Quaternary sedimentary deposits. Resultant gold placers grade upward from 1 g/m3 (0.03 troy oz/m3) and are of moderate size.

Jan 1, 1980

By William G., Pariseau

There are important phenomena in rock and soil mechanics that cannot be explained in terms of theories of homogeneous, isotropic materials. Subsidence of strata about mine openings is an example. In-situ stress measurements in nonisotropic sedimentary, and metamorphic rocks is another. Patterns of joints, faults, fractures, and similar geologic features impart directional mechanical properties to rock masses that are locally isotropic or anisotropic. On the geologic scale, the non-homogeneity of composite rock masses may, in certain instances, be dealt with as homogeneous anisotropy. The mechanics of large rock masses can, therefore, be expected to involve considerations of anisotropy. The source of the anisotropy is not relevant. Once the scale of observation is selected and the material is defined accordingly, the analysis of stress and deformation proceeds without regard to the numerical values of the material constants. There are also important phenomena in rock and soil mechanics that cannot be explained in terms of linear elasticity, theory. For the mechanical description of anisotropic geologic materials capable of de- forming beyond the limit of purely elastic strains, a plasticity theory is required.

Jan 1, 1972

By Charles H. Foreman

IN JUNE 1921, the Sullivan &lining Co., owned jointly by the Bunker Hill and Sullivan Mining & Concentrating Co. and the Hecla Mining Co, started work on the development of the Star Mine. The development work contemplated called for a crosscut from the Hecla workings on the 2000 level to the Star ledge. The length of crosscut necessary to encounter the Star ledge was approximately 8900 ft. from the Hecla shaft. By using workings formerly made by the Hecla Mining Co., only about 8100 ft. of actual crosscutting was necessary. This crosscut has a direction of S 20° W and would have a tendency to crosscut any ledges in the intervening ground.

Jan 1, 1924

By Mark Sheppard

DuRing the winter of 1937, the writer visited a West Virginia stone quarry at which the overburden is stripped hydraulically. The quarry is in a bed of limestone, about 200 ft. thick, which outcrops on a hillside above a small river; the quarry face is 175 ft. high. The top of the stone is cut in all directions by erosion channels, which are filled with a tough, red clay. Some of the larger channels are 30 ft. deep and 8 ft. wide. The thickness of the clay overburden ranges from a few inches to 10 ft., the average being estimated as less than 3 feet. The topography of the district is rugged, and the hillside above the quarry face is not suited to the use of the usual mechanical methods of moving dirt. Until hydraulicking was started 10 years ago, the overburden was dug by hand labor, dumped to the quarry floor and transported to the dump in quarry cars. Cleaning deep channels was a slow and expensive process. When hydraulicking was started, the top of the quarry face was over 200 ft. above the level of the river. The pipe line was carried in a semicircle around the top of the quarry about 200 ft. from the face and about 50 ft. above it. As there is no adequate supply of water above the quarry, the water must be pumped from the river. As quarrying progressed the pipe line was moved until the discharge points are now over 300 ft. above the river. Water is pumped from the river by a four-stage centrifugal pump having a 5-in. intake and a 4-in. discharge. It is direct-connected to a 200-hp., 2300-volt, 1760-r.p.m. induction motor. At 1760 r.p.m. the rated capacity of the pump is 750 gal. per minute under a head of 600 ft. The pump is installed in a house of frame construction on the river bank, about 10 ft. above normal water level, which also houses the pump for the mill-water supply. Above the pumps are hoisting bays into which the motors are raised whenever the river reaches flood stage. An 8-in. suction line, equipped with a strainer and a foot valve, takes the water from a concrete sump at the edge of the river. The pipe line from the pump consists of 1750 ft. of 8-in. spiral steel flanged pipe and

Jan 1, 1938

By Mark Sheppard

DuRing the winter of 1937, the writer visited a West Virginia stone quarry at which the overburden is stripped hydraulically. The quarry is in a bed of limestone, about 200 ft. thick, which outcrops on a hillside above a small river; the quarry face is 175 ft. high. The top of the stone is cut in all directions by erosion channels, which are filled with a tough, red clay. Some of the larger channels are 30 ft. deep and 8 ft. wide. The thickness of the clay overburden ranges from a few inches to 10 ft., the average being estimated as less than 3 feet. The topography of the district is rugged, and the hillside above the quarry face is not suited to the use of the usual mechanical methods of moving dirt. Until hydraulicking was started 10 years ago, the overburden was dug by hand labor, dumped to the quarry floor and transported to the dump in quarry cars. Cleaning deep channels was a slow and expensive process. When hydraulicking was started, the top of the quarry face was over 200 ft. above the level of the river. The pipe line was carried in a semicircle around the top of the quarry about 200 ft. from the face and about 50 ft. above it. As there is no adequate supply of water above the quarry, the water must be pumped from the river. As quarrying progressed the pipe line was moved until the discharge points are now over 300 ft. above the river. Water is pumped from the river by a four-stage centrifugal pump having a 5-in. intake and a 4-in. discharge. It is direct-connected to a 200-hp., 2300-volt, 1760-r.p.m. induction motor. At 1760 r.p.m. the rated capacity of the pump is 750 gal. per minute under a head of 600 ft. The pump is installed in a house of frame construction on the river bank, about 10 ft. above normal water level, which also houses the pump for the mill-water supply. Above the pumps are hoisting bays into which the motors are raised whenever the river reaches flood stage. An 8-in. suction line, equipped with a strainer and a foot valve, takes the water from a concrete sump at the edge of the river. The pipe line from the pump consists of 1750 ft. of 8-in. spiral steel flanged pipe and

Jan 1, 1938

By Mark Sheppard

DURING the winter of 1937, the writer visited a West Virginia stone quarry at which the overburden is stripped hydraulically. The quarry is in a bed of limestone, about 200 ft. thick, which outcrops on a hillside above a small river; the quarry face is 175 ft. high. The top of the stone is cut in all directions by erosion channels, which are filled with a tough, red clay. Some of the larger channels are 30 ft. deep and 8 ft. wide. The thickness of the clay overburden ranges from a few inches to 10 ft., the average being estimated as less than 3 feet. The topography of the district is rugged, and the hillside above the quarry face is not suited to the use of the usual mechanical methods of moving dirt. Until hydraulicking was started 10 years ago, the over- burden was dug by hand labor, dumped to the quarry floor and transported to the dump in quarry cars. Cleaning deep channels was a slow and expensive process. When hydraulicking was started, the top of the quarry face was over 200 ft. above the level of the river. The pipe line was carried in a semi- circle around the top of the quarry about 200 ft. from the face and about 50 ft. above it. As there is no adequate supply of water above the quarry, the water must be pumped from the river. As quarrying progressed the pipe line was moved until the discharge points are now over 300 ft. above the river. Water is pumped from the river by a four-stage centrifugal pump having a 5-in. intake and a 4-in. discharge. It is direct-connected to a 200-hp., 2300-volt, 1760-r.p.m. induction motor. At 1760 r.p.m. the rated capacity of the pump is 750 gal. per minute under a head of 600 ft. The pump is installed in a house of frame construction on the river bank, about 10 ft. above normal water level, which also houses the pump for the mill-water supply. Above the pumps are hoisting bays into which the motors are raised whenever the river reaches flood stage. An 8-in. suction line, equipped with a strainer and a foot valve, takes the water from a concrete sump at the edge of the river. The pipe line from the pump consists of 1750 ft. of 8-in. spiral steel flanged pipe and 800 ft. of 4-in. spiral steel flanged pipe. The connection from the pump

Jan 1, 1938

The Candelaria deposit was first discovered by Spanish prospectors in 1863; but its silver ore was not exploited on a large scale until 1873, when the Northern Belle mine was opened. The first Northern Belle mill was built at Belleville, 13 km (8 miles) to the northwest. From 1876 to 1886 the Northern Belle paid over $1 million in dividends each year. In 1882 a 43 km (27 miles) pipeline was built to Candelaria from the White Mountains to the west to furnish water for the mills and town at Candelaria; it was 100 mm (4 in.) in diameter, hand- riveted, and made in England. Sections of this pipe are used. as claim posts today. At its terminus at Metallic City, above Candelaria, a large concrete reservoir was built and called Metallic Lake. It is being used today by Candelaria Partners. Production at Candelaria declined after the mid-1980s, revived for a short period in the early 1890s, and during the 1920s a 300 ton mill operated for three years. Candelaria Partners is owned jointly by Occidental Minerals and Congdon and Carey, with Occidental being the operator. An engineering study done in mid-1979 resulted in engineering for the plant commencing in late that year and heap leaching being started in September 1980. Two million tons of ore per year containing 108 g Ag and 0.28 g Au per mt (3.15 oz Ag and 0.008 oz Au per st) will be mined from the Lucky Hill and Mount Diablo .areas about 1.6 km (1 mile) up Pickhandle gulch from the heap leaching operation. Waste -to -ore ratio is about 2:l. Ore is mined by front-end loaders and hauled by 85 ton trucks to the primary crusher, a 1,070 by 1,780 mm (42 by 70 in. ) gyratory. Primary crushed ore is conveyed to a stockpile of 10,000 tons live capacity from which it is reclaimed by hydraulic reciprocating feeders and conveyed to the crushing and agglomeration facility. This unit is composed of a 1.8 m by 2.6. m (6 ft by 12 ft) double deck screen, a 2.1 m (7 ft) standard cone crusher, and three conveyors. At the conveyor transfer points NaOH and NaCN solution is sprayed on the ore as it falls through a row of 75 mm (3 in.) hanging steel bars. During this process of turning and dividing the ore, the fines agglomerate with the coarse ore to give better percolation for the leach solution on the heaps. Provision was made for the later addition of 2 kg per mt (4 lb per st) of cement to aid in the agglomeration. Agglomerated ore at minus 50 mm (2 in.) size is then conveyed to a 390 mt (430 st) loadout bin where it is transferred to 85 ton rear dump trucks for haulage to the leach pads.

Jan 1, 1981

By Linn Bradley, W. W. Strong, H. D. Egbert

A hot-dry method of cleaning the gas from blast furnaces has been shown1 to conserve the sensible heat energy of the gas, and in general it thus permits of a higher flame temperature. The electrical method of cleaning gases removes from 98 to 99 per cent. of the suspended matter in one operation when operating properly, thus making possible the use of a very clean gas for burning in the stoves. The use of such clean gas thus makes feasible a different type of hot-blast stove construction from that being used at present. It therefore seems advisable to construct the stoves from the viewpoint of getting the maximum amount of heat energy into the products of combustion. By preheating the air added to the top gas and obtaining a better mixture of the air and top gas, less excess air will be required and a higher flame temperature obtained from the combustion of the top gas. The interior of the stove should be provided with a kind of brick that will withstand a high flame temperature, will conduct the heat energy along certain directions, insulate the flow of heat along other directions and at the same time store heat energy so as to make the flow of such energy more uniform. It is quite likely that these results can be best obtained by using several kinds of bricks: one type which is a good insulator of heat energy to be used for lining the stove and for making partitions through the stove to restrain the flow of heat; another type of brick or material with high refractory qualities; and a third type which will be a good conductor of heat energy. The surface of the stove exposed to the atmosphere should be a minimum in comparison to the mass of checkerwork which it is possible to obtain, due regard being paid to other conditions. With clean and hot furnace top gas it will be advisable to consider a rearrangement of the checker work in the stoves. One season for the present size and shape of checker openings is to provide facility for

Jan 1, 1917

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. Howell Clevenger, Bhupendranath Ray

Formerly great divergence of opinion existed in regard to the influence of copper in steel, as affecting its various physical properties. More recently the investigations of Stead,l Breuil,2 Wigham,3 Burgess and Aston: Ball and Wingham,5 Miiller; Dillner,7 and others have shown fairly good agreement in the major points. Copper steel has been shown by these investigators to possess special properties which make it valuable for certain uses. The general range of composition of the useful alloys of copper and iron, and of copper, iron, and carbon, has been quite definitely established, for the former by Burgess and Aston,8 and for the latter by Breuil9 and Stead.10 Stead11 has removed one difficulty in reconciling certain of the earlier data by calling particular attention to the influence 'of carbon in copper steels. Another point which caused confusion in the earlier investigations, particularly in regard to " red-shortness," was the presence of sulphur. Failure to note the presence of sulphur often led to attributing " red-shortness" to the presence of small amounts of copper. Turner12 has called attention to the fact that where copper exists in an iron ore it often occurs as the sulphide, so that the presence of copper in steel made from such ores might also, necessarily, mean the presence of sulphur. It is evident that the effect in this case might be quite different than where pure copper was added.

Jan 1, 1914

By J. Stubbins

1 1.4-1. Source of Water. Precipitation in the form of rain or snow is the original source of water that enters most surface mines. The precipitation, which does not escape to the atmosphere, either becomes surface run- off or infiltrates into the ground to become ground water (3,4). Evaporation and surface run-off can be measured. The water that infiltrates the ground disappears from view and is more difficult to evaluate. In general, it moves in a pattern similar to the surface run-off, but at a much slower rate, since its progress is impeded by many subsurface conditions. The infiltrating water descends through the soil and rock by whatever passageways permit flow until the water table or top of the zone of saturation is reached. Though usually pictured as a level surface, it is dependent upon the topography and the permeability of the enclosing material, and therefore it can be as erratic and unpredictable as the subsurface geology. Ground water may become surface run-off again at a lower point, and likewise, surface run-off of lakes and streams can infiltrate to become ground water. Should the ground water be confined to a permeable zone overlain by an impermeable barrier, a pressure is created. Such water is said to be artesian if the pressure head is sufficient to raise the confined water to the surface. These conditions are more commonly found in relatively flat-lying bedded formations. 11.4-2. Recognition and Estimation of Water. Dewatering and flood control should be considered by the design engineer at an early stage in all surface mine planning. Recognition of the possible problems, and making allowance for their solution, is an inexpensive precaution. Circumstances which should lead to investigation are: 1. High rainfall 2. Experience of other mines in the area 3. Evidence of water and caving in drill holes 4. Presence of springs and artesian wells 5. Geological evidence of aquifers.

Jan 1, 1968

By R. B. Walker, R. Palmer, L. C. Bryner

The purpose of this investigation was to further study some of the factors affecting the air oxidation of sulfide minerals. Data obtained from laboratory studies on the effect of temperature, concentration of iron salts, pH, and other catalysts on the biological and nonbiological oxidation of sulfide minerals are presented. The oxidation of copper sulfide minerals and subsequent leaching of the soluble copper has long been a profitable operation. The chief oxidizing agent is oxygen from the air. This reaction is influenced by a number of different factors such as temperature, light, hydrogen ion concentration, oxidizing agents, and various catalysts. Sullivan1 found that copper sulfide minerals are slowly oxidized by the air and that dilute sulfuric acid and iron (111) increased their solubility. At that time it was not known that bacteria were involved in the oxidation of the sulfide minerals. It has been shown that pyrite, molybdenite, free sulfur, and the copper sulfide minerals found in the waste rock dumps in Bingham Canyon, Utah, are largely oxidized by chemosynthetic autotrophic microorganisms found in the leaching streams.2-6 Similar organisms have been isolated in other mining areas7,8 and from the effluents from coal rnines.9 The following are some of the questions that have arisen regarding the mechanisms of the reactions involved in the leaching process. How dependent is the leaching operation upon the biological reactions? How does the temperature effect the biological and non-biological reactions? How does the concentration of iron (II) and (111) effect the rate of oxidation of copper (II) sulfide? Does the iron act as a catalyst? Does activated carbon act as a catalyst on the oxidation of sulfide minerals? What effect does the hydrogen ion concentration (pH) have on the above reactions? The purpose of this paper is to report the results of laboratory studies on these problems. EXPERIMENTAL The apparatus used in this investigation was an air lift percolator described in our earlier work.' The percolator consisted essentially of an air lift which enabled the leaching solution and microorganisms to pass freely through a suspension of mineral (substrate) in Ottawa sand. The air lift served the additional purpose of maintaining an ample concentration of oxygen and carbon dioxide in the nutrient (leaching) solution. (See Fig. 1.) The percolators were so constructed that they could be immersed in liquid constant temperature baths for temperature studies. The thermostats were controlled to +0.5°C. The reactions were all run in the dark. The percolators were charged with one to three g of mineral, dispersed in 100 g of Ottawa sand and 100 ml of nutrient or leaching solution. The sand was purified to remove traces of iron by digesting with concentrated hydrochloric acid and washng with distilled water until free from chloride. After drying, the sand was ready for use. The compressed air supply was regulated to 5 psi and bubbled through an aerator containing distilled water, which was maintained at the same temperature as the bath used. The air was filtered through cotton plugs to maintain sterile conditions. The tops of the

Jan 1, 1968