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By K. Harms, T. D. Jones

TO unload large tonnages of lead bullion cast in 100-lb. bars is a problem which has confronted the lead refineries for many years. The bars, on arrival, must be restacked for unloading by truck or hand-charged to charging machines. Selected crews of labor, capable of working at sustained speeds for five- or six-hour periods, are required for this exhausting task. Due to the fatigue involved crew personnel changes rapidly. During the war it became increasingly difficult to maintain crews, especially in critical labor areas, so that the development of casting and shipping ten-ton lead blocks from the smelters to the refineries became a must in our operating program. For many years lead refineries, when making interdepartment or intradepartment transfers, have cast lead blocks weighing five or ten tons which are easily handled by crane by merely attaching the crane hook to a steel anchor embedded in the frozen lead blocks. In previous years, the practice was to use regular smelter slag pots which, due to their rounded shape, produced what were called "buttons" by the refinery personnel. The button-type blocks were unsafe as they rolled and tipped easily and were unsatisfactory for stacking since no piling pattern could be followed. To remedy this situation, flat-bottom molds came into general use and later the bottoms were made concave to prevent slipping.

Jan 1, 1946

SIDNEY ROLLE, Chrome, N. J.-I should like to ask if Mr. Bach does not think it rather harmful to let the amen depend entirely on the company, whether it would not be a good plan to let the men aid a little in the support of these organizations? Also whether he has had any experience with utilizing the Y. M. C. A. in this connection. E. E. BACH, Ellsworth, Pa.-That question is always raised whenever there is discussion upon this subject. So far as we are concerned that would not help a bit; our people do not have that feeling, because .we have a company which these men recognize has a corps of very sympathetic officials in every way. I remember the assistant general manager lived out in one of these mining towns for a while. His difficulty was that every man, no matter what, who or where he was, if he had, or thought lie had, a, complaint, had the right of way to his office, and didn't have to stop at all. That is the spirit that predominates all through it. This is also the spirit of the mine foremen, two of whom were trained in our own night school, and it is the spirit of the fire boss and the general spirit of the people. They do not feel that we are trying to feed them with a spoon. As to the Y. M. C. A., we feel that it would be impracticable for us. R H. VAIL, New York, N. Y.-It would be very interesting if Mr. Bach would tell a little more about the control of the dispensing of intoxicants. There are many communities where the state intends to do that but does not succeed very well, and as Mr. Bach has spoken of the speakeasies on the environment of these towns, it would he well to know more about the problem of eliminating those.

Jan 4, 1918

By H. C. Weed

THE Inspiration Consolidated Copper Co., located in the Globe-Miami district at Inspiration, Ariz., became a producer of copper in 1915. From 1915 until 1948, 116,278,000 tons of ore were produced from underground workings by the block-caving system of mining. In April 1948, open-pit ore production from portions of the ore body was started and earlier plans for block-caving in these areas were abandoned. The change was due to economic conditions.

Jan 8, 1950

By H. Rumfelt

The author demonstrates an approach for analyzing overcasting requirements for a stripping project. This approach to the problem employs indicated trends in the relationship of the weight of the machine (only electropowered draglines are considered) to its ability to do stripping work. Strip mining in the domestic coal industry is contributing a greater proportion of overall production than at any time in the past. U.S. Bureau of Mines statistics show that the percentage of total production mined in 1958 by stripping methods increased to more than 30 pct, up from 22 pct in 1951. The tendency to favor strip and open pit mining exists in other mining activities, but such trends are not as clearly outlined as in the coal industry. Significantly, the depth of overburden in coal properties considered suitable for stripping is also increasing. A depth of 50 to 60 ft of overburden was once considered to be about the maximum amount that could be handled economically. Today, a depth of 80 to 100 ft is frequently taken into consideration for simple overcasting operations (i.e., excavating overburden and dumping it in the spoil position where it remains indefinitely). Accompanying these changes, there have been continuing studies searching for procedures to appraise deeper stripping problems. Although the only known practicable way to accurately evaluate a proposed stripping venture is to make studies (including cost estimates) step by step, it is nevertheless desirable to have a quick method for making preliminary evaluations of potential strip mining operations. A recently devised digital computer method of providing a preliminary simple overcasting analysis for a stripping prospect relates the indicated trends in the relationship of the weight of the machine to its ability to do stripping work. In addition, this approach employs situations where the geometry of each cut-and-spoil section is assumed to take certain defined relationships for varying overburden depths. Slopes are necessarily considered to be stable, which means such practical factors as the mechanics of soils are neglected for the sake of convenience. Definition of Terms: Photographs A and B show a shovel-type and a dragline-type of simple overcast operation, plus terms used to denote certain parts of a pit. Pits that follow a straight line when projected on a horizontal plane are referred to as straightaway. It has been assumed that all section drawings accompanying this article pertain to straightaway cuts which, in turn, means that areas can be compared by relative volume. The volume of overburden is measured in virgin cubic yards. Material displaced from its virgin state to a spoil area normally occupies a larger volume than when in situ; the difference is termed swell, and it is expressed as a percentage of the original volume. For example, if the original cut volume is denoted V cu yd, and the spoil volume is 1.2 V cu yd, the swell is a positive 20 pct. MAXIMUM USEFULNESS FACTOR (MUF) CONCEPT Men experienced in overburden removal give emphasis to the weight of the machine in relation to its value as an excavator. From the writer's understanding, value has been only loosely defined and discussed generally and vaguely. The concept is considered important, however, since it is based upon the observations of many men. It provides a clue for the approach taken to solve deep stripping problems. In the procedure described below, value has been arbitrarily established as the product of the nominal dipper size of the shovel (or dragline) times a functional dumping reach. MUFs—Shovel: A shovel's capability to handle thick overburden is usually limited by its ability to dispose of the spoil. Thus the dumping reach, as well as its dumping height, is significant. The angle of repose of spoil material must also be considered. Although such slope angles vary among different mines, jobs, and materials, a slope of 1.25 to 1, frequently found in both planning and practice, is used by this author in the illustration of this analytical method. The relationship of the shovel's geometry to a stripping operation is shown in Fig. 1. To gain maximum advantage in constructing spoil piles, the shovel is placed so that its tracks adjacent to the spoil are as close as possible to the rib of the coal deposit. A vertical plane passing through the line

Jan 1, 1961

By John Conley

THE production of lime by the burning or calcination of limestone, including all varieties from true dolomites and magnesian limestones to high-calcium types, continues as one of the essential basic industries. In 1937, a total of 4,124,165 tons of lime valued at approximately $30,091,000 was produced in the United States.1 The importance of any contribution to our knowledge on the mechanism of the lime-burning process is, therefore, obvious. The burning of limestone has been practiced since the earliest days of our civilization, and the literature on the various phases of the industry is voluminous. Calcination of the various types of limestone has been studied by many previous investigators, but some points require addi-tional research. Each occurrence of limestone provides an opportunity for specific investigation, because the characteristics of stones in different localities vary. Knibbs13 has given an excellent discussion of available data on the theory of lime and magnesia burning. The thermal dissociation of calcium carbonate has been studied more recently by Johnston,2 Smyth and Adams,3 Ralston et al.,4 Furnas,5 Potapenko,6 and Southard and Royster.7 As a result of these investigations the dissociation temperature at atmospheric pressure has been established as 894° to 898° C. The dissociation of dolomite and magnesite, although studied by several investigators, 4,6,8 still requires clarification by additional data. The calcination of limestone was studied for a twofold purpose: (1) to extend our information on the mechanism of the lime-burning process and (2) to standardize the conditions for calcination of high-calcium and dolomitic limestones for subsequent hydration tests on the lime.

Jan 1, 1939

By E. Teghtsoonian, A. R. Causey

THE fact that the surface plays an important role in the plastic deformation of a single crystal has been demonstrated by several investigators. The surface conditions have been altered using oxide films,1'2 alloyed metal layers,3, 4 deposited metal films,' and polycrystalline layers.677° The resulting effects on the initiation and subsequent plastic deformation have been investigated extensively except for the case of a single crystal surrounded by a polycrystalline layer of the same material. The purpose of this paper is to describe an investigation of the effect of a polycrystalline surface layer on the yield stress, the critical resolved shear stress, and the work-hardening slope of tin single crystals. The tin single crystals used in this study were grown from the melt using material having an initial purity of approximately 99.999 + pet with lead and iron the major impurities. Fig. 1 shows the orientation of the (110)[001] slip system with respect to the crystal surfaces. The crystals had a 1/4 by 1/4 in. cross section and a gage length of 2 in. The polycrystalline surface layer was produced by abrading the single-crystal specimens with emery paper. Fig. 2 shows the broken Debye rings obtained from the deformed surface. The thickness of the layer was controlled by the number of passes on different grades of paper and was determined by electropolishing in dilute HC1, until the grains produced by the abrasion had all been removed. Metal-lographic examination showed that the grain size varied with depth, increasing from 10 µ right at the surface to 40 µ in the interior of the layer. The yield stress, critical resolved shear stress,

Jan 1, 1965

By John E. Conley

The production of lime by the burning or calcination of limestone, including all varieties from true dolomites and magnesian limestones to high-calcium types, continues as one of the essential basic industries. In 1937, a total of 4,124,165 tons of lime valued at approximately $30,091,000 was produced in the United States.' The importance of any contribution to our knowledge on the mechanism of the lime-burning process is, therefore, obvious. The burning of limestone has been practiced since the earliest days of our civilization, and the literature on the various phases of the industry is voluminous. Calcination of the various types of limestone has been studied by many previous investigators, but some points require additional research. Each occurrence of limestone provides an opportunity for specific investigation, because the characteristics of stones in different localities vary. Knibbs13 has given an excellent discussion of available data on the theory of lime and magnesia burning. The thermal dissociation of calcium carbonate has been studied more recently by Johnstonr2 Smyth and Adams,3 Ralston et al.,4 Furnas,6 Potapenko,6 and Southard and Royster.7 As a result of these investigations the dissociation temperature at atmospheric pressure has been established as 894" to 898' C. The dissociation of dolomite and magnesite, although studied by several investigators,4,6,8 still requires clarification by additional data. The calcination of limestone was studied for a twofold purpose: (I) to extend our information on the mechanism of the lime-burning process and (2) to standardize the conditions for calcination of high-calcium and dolomitic limestones for subsequent hydration tests on the lime.

Jan 1, 1942

By John E. Conley

The production of lime by the burning or calcination of limestone, including all varieties from true dolomites and magnesian limestones to high-calcium types, continues as one of the essential basic industries. In 1937, a total of 4,124,165 tons of lime valued at approximately $30,091,000 was produced in the United States.' The importance of any contribution to our knowledge on the mechanism of the lime-burning process is, therefore, obvious. The burning of limestone has been practiced since the earliest days of our civilization, and the literature on the various phases of the industry is voluminous. Calcination of the various types of limestone has been studied by many previous investigators, but some points require additional research. Each occurrence of limestone provides an opportunity for specific investigation, because the characteristics of stones in different localities vary. Knibbs13 has given an excellent discussion of available data on the theory of lime and magnesia burning. The thermal dissociation of calcium carbonate has been studied more recently by Johnstonr2 Smyth and Adams,3 Ralston et al.,4 Furnas,6 Potapenko,6 and Southard and Royster.7 As a result of these investigations the dissociation temperature at atmospheric pressure has been established as 894" to 898' C. The dissociation of dolomite and magnesite, although studied by several investigators,4,6,8 still requires clarification by additional data. The calcination of limestone was studied for a twofold purpose: (I) to extend our information on the mechanism of the lime-burning process and (2) to standardize the conditions for calcination of high-calcium and dolomitic limestones for subsequent hydration tests on the lime.

Jan 1, 1942

By John E. Conley

THE production of lime by the burning or calcination of limestone, including all varieties from true dolomites and magnesian limestones to high-calcium types, continues as one of the essential basic industries. In 1937, a total of 4,124,165 tons of lime valued at approximately $30,091,000 was produced in the United States.1 The importance of any contribution to our knowledge on the mechanism of the lime-burning process is, therefore, obvious. The burning of limestone has been practiced since the earliest days of our civilization, and the literature on the various phases of the industry is voluminous. Calcination of the various types of limestone has been studied by many previous investigators, but some points require additional research. Each occurrence of limestone provides an opportunity for specific investigation, because the characteristics of stones in different localities vary Knibbs13 has given an excellent discussion of available data on the theory of lime and magnesia burning. The thermal dissociation of calcium carbonate has been studied more recently by Johnston,2 Smyth and Adams,3 Ralston et al.,4 Furnas,5 Potapenko,6 and Southard and Royster.7 As a result of these investigations the dissociation temperature at atmospheric pressure has been established as 894º to 898º C. The dissociation of dolomite and magnesite, although studied by several investigator,4, 6, 8 still requires clarification by additional data. The calcination of limestone was studied for a twofold purpose: (1) to extend our information on the mechanism of the lime-burning process and (2) to standardize the conditions for calcination of high-calcium and dolomitic limestones for subsequent hydration tests on the lime.

Jan 1, 1939

By George W. Mitchell

THE property of Eureka Corp. Ltd. is located in the approximate geographic center of Nevada, 2 miles from Eureka, the county seat. The great sources of power, the Colorado, Snake, and Salmon Rivers and the rivers of northern California, are 300 to 500 miles distant, and no lines serve areas closer than 150 miles. Fuel for diesel and steam generation is available in Utah, 300 to 400 miles to the east. Eureka's railhead is 80 miles north where two trunk lines cross the county. A spur line serves Ely, 77 miles east. Good highways connect Eureka to the railheads. Activity in the Eureka mining district began in the early 1870's. The oxidized high grade lead-sil-ver-gold ore terminated against the footwall of the Ruby Hill fault, and in 1890 the main operations ceased. In 1938 Eureka Corp. Ltd. discovered ore in the hanging wall of the fault by diamond drilling. The history of Eureka in the late 1800's indicates that there was some water at 600 to 800 ft in the old workings, probably accumulations above the water table which did not seriously interfere with mining operations. Both the Locan and Richmond shafts were sunk to a level below the table, but apparently the only serious difficulty with water occurred in the Locan. The steam pump used when the last work was done on the 1200 level in 1923, many years after exhaustion of the main orebodies, is still installed on the Locan 900 level. The capacity was about 500 gpm, lifting 750 ft to the 100 level, which connected with the surface. In addition to this. bailers were used to keep the 1200 level free of water. It is said that pumping in 1923 lowered the water in the Holly shaft, about a mile and a half away, but this seems doubtful. The pumping was of short duration because no ore was found. When work at the new Fad shaft was started in 1941 Eureka Corp. Ltd. engineers were fully aware of the probability of encountering water in large volume. Their primary exploration and development had to be carried on at the 2250 level. The first water was encountered at 300 ft. This was undoubtedly surface drainage in the bedding of the Pogonip limestone and was less than 100 gpm. The fractured, loose Hamburg dolomite at the water table was not well cemented, and relatively little water, 300 gpm, percolated through it with difficulty. At 1350 ft well-cemented dolomite containing some open fractures was encountered. These fractures produced the first water of consequence, 750 gpm. At 1700 ft the volume was 1000 gpm increasing to the maximum during shaft sinking, 1600 gpm, at the 2000 level. Secret Canyon shale, a dry formation, was entered at 2100 ft, where a concrete water ring was placed to catch all of the water. The volume decreased rapidly to a constant flow of 1200 gpm. Below 2100 ft the shaft and stations remained in the shale and water was not a problem. Several faults of moderate displacement, including the reverse Martin fault, had been intersected during the traversing of 1000 ft of wet Hamburg, but no undue quantities of water were encountered. Observations in the diamond drill holes in the ore zone area showed a rapid lowering of the water table. The shaft was flooded when it left the dry shale and entered the water-bearing Eldorado dolomite on the 2250 level, crossing a fissure which paralleled the Martin fault. High pressure water doubled the volume then being pumped. Pipe failure through a water door bulkhead was a contributing factor. Immediately following this flooding in March 1948 preparations were made to recover the shaft as rapidly as possible by increasing power and pump capacities as needed. Measurements before flooding indicated the water could be lowered at a fast rate. However, the water table did not recede as rapidly as expected and volumes required to lower the water in the shaft were higher. Obviously the size of the main water channel on the 2250 level was increasing because of erosion, allowing greater volumes to enter the workings and draining beyond the cone originally being drained during shaft sinking. Eroded material was being deposited in the shaft below the 2250 level in serious proportions. In December 1948 a second flooding of the Fad shaft was allowed for the purpose of reassessing existing conditions and studying alternate methods of attack. The detailed geology of the Eureka mining district, see Fig. 1, has been described during the past 75 years by many geologists.' Only the general features and those which seem to affect the drainage problem will be discussed. The old ore zone, mined between 1870 and 1890, is located in a wedge-shaped block of Eldorado dolomite between the footwall of the Ruby Hill fault and the underlying Prospect Mountain quartzite, see Fig. 2. Production of high grade oxidized lead ore containing high values in gold and silver has been variously estimated at $50 to $90 million. The tonnage mined was probably close to 1,500,000, nearly all of which was found above the water table. The new ore, discovered by diamond drilling in the hanging wall of the Ruby Hill fault, is a flat-

Jan 1, 1954

By J. H. Frye, J. L. Scott, J. W. Woods

THERE are at least two effects of temperature on the stress required to produce plastic flow. 1) Metals plastically strained for the same amount at different temperatures possess different stress-strain curves and different internal structures (e.g., nature and density of imperfections). 2) There is both experimental and theoretical evidence that the stress necessary to produce plastic flow may vary with temperature even though the internal structure remains constant.1-5 Much of the data in the literature is difficult to interpret because these two effects are simultaneously present. In the work reported here, they are considered separately. Standard 0.505 in. tensile specimens were prepared from high-purity Mond nickel and OFHC copper. Nickel specimens were annealed in hydrogen at 1150°C for 30 min and furnace cooled. This produced an average grain diameter of 1.4 mm. Copper specimens were annealed in vacuum at 525°C for 6 hr and furnace cooled. This produced an average grain diameter of 0.11 mm. Nickel tensile specimens were prestrained in tension at 500°K, whereas copper tensile specimens were prestrained in tension at room temperature. All specimens were cooled to 74°K and heated to room temperature. After these treatments all of the tensile specimens of a given set possessed identical structures. All specimens were returned to the tensile machine and tested at various temperatures and constant cross-head speed. Many of the tensile curves showed well defined yields but, in order to minimize personal judgment, the yield was always chosen as the stress corresponding to 0.2 pct plastic deformation. Results are summarized in the graph of yield vs temperature of test, Fig. 1. All of the specimens corresponding to any one line have had identical treatments and, therefore, have identical internal structures. Thus, each line gives the relation between yield stress and temperature for one constant internal structure. The lowest line, which is for annealed nickel, shows no measurable variation of yield with temperature. This is similar to Blewitt's6 results on copper single crystals, but a variation in yield with temperature was found by Andrade and Henderson' in single crystals of nickel, and by Car-recker and Hibbard8 in annealed polycrystalline copper. The three other lines, which are for prestrained materials, show that there is a linear relation between yield stress and temperature when the internal structure remains constant. Such a relation might be expected from a rate equation such as that of Kauzmannl on the assump- tion that the activation energy is either independent of or linear with temperature. These results would also be expected if the shear moduli of copper and nickel are linear functions of temperature for, as Cottrell and Stokes have pointed out," in all dislocation theories of strength the flow stress appears proportional to the shear modulus of the material. In copper and nickel, since the modulus of elasticity decreases roughly linearly with increase in temperature," it is probable that the modulus of shear also decreases approximately linearly with increasing temperature. A detailed discussion of the effect of these and other factors on the temperature dependence of flow stress in face-centered-cubic metals has recently been given by Seeger.4

Jan 1, 1958

By T. D. Yensen, R. K. McGeary

While we have not been primarily interested in the determination of hydrogen in the alloys that we have been dealing with, we arc very glad to cooperate in this symposium on sampling and analysis for hydrogen, as we can never tell when we may find it desirable to perform such analyses. Vacuum-fusion Method Our primary interest has been oxygen and carbon, in the analysis of which we think we have become very proficient. Our oxygen-analyzing apparatus is of the vacuum-fusion type (Fig. I). This employs Dr. Derge's type of graphite cruciblc and slit sleeve, with graphite powder between, for heat insulation (Fig. 2) supported in a clear quartz bulb connected to the Pyrex glass system by means of a carefully made ground joint. The latter is kept cool by means of an air blast from a circular copper tube surrounding the joint. The crucible is heated by high-frequency induction and the 24 one-gram samples successively dropped in from the horizontal tube at the top by means of a magnet. The gases (H², CO, N2) are passed through the CuO catalyst held at 300°C. (converting CO to CO²I, H² to H²O) and the CO² and H2O frozen out in the main liquid air trap, while N2 is evacuated. At the end of the operation, shown by the termocouple vacuum gauge indicating a pressure of 10-4 mm. Hg, the analyzing system is isolated by means of the mercury cutoffs, and the CO² and H2oare determined by transferring first the CO2 and then the HzO to the small trap at the right. This is done by placing liquid air on the small trap, then removing the liquid air from the large trap for one minute, which is sufficient to evaporate the Co2 and transfer it to the small trap. The CO2 is then measured by closing the mercury cutoff at the extreme right, removing the liquid air from the small trap, warming it up to room temperature, adjusting the mercury level to one of the three marks on the glass tube (corresponding to three volumes calibrated so that I mm. Hg of CO² corresponds to 0.001, 0.002 and 0.010 mg., respectively, of oxygen) and finally reading the pressure. Liquid air is now replaced on the large trap, the mercury cutoff at the right opened, transferring the COz back to the large trap. The cutoff to the pumps is then opened, the liquid air removed for one minute, thus evacuating the CO², the pumps again cut off and the H2 O transferred to the small trap as before, except that it is now necessary to warm the large trap with a Bunsen burner in order to evaporate all the H2O. This requires about three minutes, as the inner tube must be heated to at least 100°C. In order to prevent H²O from being adsorbed on the glass walls, the glass tubing beyond the catalyst (i.e., to the right in Figs. I and 3) is kept heated to about 150°C. by means of Nichrome ribbon as shown in Fig. 3. For a I-gram sample the above amounts of oxygen (0.001, 0.002 and 0.010 mg.) 35

Jan 1, 1945

By J. P. Hirth, G. M. Pound

The flux of evaporation J from {loo) , {111), and (110) planes of silver single crystals and from poly crystalline specimens was measured at 744°C under a high vacuum using a time average weight loss method. The evaporation coefficient a was computed from the measured flux and vapor pressure data through the relationship J = aPe/21rmkT. The evaporation coefficients for the close packed planes and the poly crystalline specimens were found to be about 0.4 providing tentative confirmation of the theoretical limiting law for evaporation of metal crystals, a = 73(P/Pe) + % . The (110) planes faceted to form (111) planes, resulting in a greater evaporating area and larger a . The evaporation morphology of the plane surfaces and the surfaces of spherical single crystals is described in some detail. RECENTLY a detailed theoretical analysis of the mechanism and kinetics of metal crystal evaporation was made.1'2 A mechanism for vaporization involving the recession of monatomic ledges across a metal surface was deduced, and evaporation rate equations were derived by considering the dynamics of these ledges. In most metals, the process of evaporation is one of dissociation of atoms from ledges, surface diffusion, and desorption. The gross vaporization rate is controlled by the surface diffusion and desorption rates. It was shown that the monatomic ledges, under the action of surface diffusion gradients, tend to assume a terminal velocity and spacing. Further, it was shown that crystal edges and grain boundaries are a ready source for monatomic ledges. In the case of evaporation of large, perfect crystals with clean, low-index surfaces, the kinetic treatment1 leads to a limiting law for crystal evaporation: where J is the gross flux of vaporization, a is the evaporation coefficient, Pe the equilibrium vapor pressure, m the atomic mass, k Boltzmann's constant, T the absolute temperature, and P the actual pressure of vapor over the crystal surface. Imperfections such as grain boundaries, pores, cracks, and macroscopic steps should have the effect of causing positive departures from this limiting law. Such an imperfection is a source for monatomic ledges, which, near their origin, are closely spaced and thus give rise to higher evaporation rates there. However, under many experimental conditions,' this effect should be negligible due to the relatively low concentration of imperfections. In general, a more important cause of positive departure should be screw dislocations, and the effect of these has been considered in some detai1.2,3 Monatomic ledges emanate from points where the surface is intersected by a dislocation with a screw component normal to the surface. The dynamics of ledges arising from a screw dislocation are similar to those of ledges arising from a crystal edge. There is a perturbed region near the origin where the monatomic ledges are, under certain experimental conditions, closely spaced and the

Jan 1, 1960

By A. L. Mular

The flotability of crushed zinc oxide pellets which were doped to produce more n-type or less n-type (more p-type) properties was studied with a Halli-mond tube. Flotation data are presented to show that the pH at which 50% recovery occurs and the collector concentration for 50% recovery depends upon the concentration of doping agent added substitu-tionally for zinc in zinc oxide. The time of sintering of pellets has a slight effect on flotation behavior. A speculative model is proposed to explain the observed behavior in light of the fact that zinc oxide is a semiconductor and has a space charge region at the surface. Experimental verification is necessary before the proposed model can be established as correct. In recent years, the attachment of flotation reagents to minerals has been shown to depend upon the behavior of an electrical double layer which exists in the liquid near the mineral-liquid interface.' Thus, a mineral placed in water will undergo a surface re-arrangement, most likely be preferential dissociation of lattice atoms or by surface adsorp-tion-desorption reactions, to form a dissociation electrical double layer with the solution. Among the many factors which influence adsorption processes, the past history of the mineral itself is one of the most important. Because of trace impurities, minerals from different localities may exhibit marked differences in flotation behavior. This is not unexpected, since investigators who have studied the detailed structure of the double layer's3 have known for some time that the impurity content of many minerals may exert strong influence on double layer variables. For this reason, their studies were usually made with minerals of high purity. The precise relationship between impurities in a mineral and double layer parameters such as capacity and electrostatic potential has been obscure. However, impurities in a semiconductor type mineral which is in contact with a solution have been recently shown to produce changes in the properties of a space charge layer which extends into the bulk of the semiconductor to a distance of one micron or more.4 The double layer is influenced by the space charge layer, whose origin is best discussed in terms of the band theory of solids. A qualitative description of the behavior of impurities in the bulk of a semiconductor, from a chemical point of view, is useful at this point. Briefly, impurities or defects (lattice vacancies, interstitials, etc.) in semiconductors such as PbS, ZnS, FeS2, Cu2S, NiS, NiO, Fe2O3, ZnO, CdS, MnO2 and many others, behave somewhat like acids or bases in water. For example, water dissociates into hydrogen and hydroxyl ions according to

Jan 1, 1965

By F. Louis Grammer

THE difference between blast-furnace gas and ordinary producer-gas is chiefly that blast-furnace gas is richer in CO, and poorer in hydrocarbons and hydrogen, as is shown in Table I. TABLE I.-Volume-Composition of Producer-Gas and Blast-Furnace Gas.. Producer-Gas. Blast-Furnace Gas. Per Cent. Per Cent. (Volume.) (Volume.) CO . 22 to 33 25.5 00., . 6 to 3 13.0 H . . 13 to 7 1.5 CH, 3 to 1.5 0.0 N . 54 to 60 60.0 If natural gas, water-gas, or by-product coke-oven gas be available, blast-furnace gas can be made to resemble very closely in calorific value the average producer-gas by merely mixing in one of these products, which are richer in hydro-carbons and, hydrogen. Blast-furnace gas, however, possesses a virtue which is sometimes absent in producer-gas, especially with coals very high in sulphur-namely, that blast-furnace gas is relatively free from sulphur. In some regions the sulphur in coal is an acute problem, and it is highly desirable to find uses for the products which are specially high in sulphur-content. An advantage would be gained if coal was coked before charging into the gas-producers, but this adds to the expense. Sulphur may be partly eliminated from coal by washing it, the success depending upon the form in which the sulphur occurs, and upon the efficiency of the washer, but it cannot always be removed to the extent desired at most American plants. It is not practicable to remove all the sulphur from producer-gas by washing it, though Mr. Reinhardt, of Germany,

Sep 1, 1908

By M. E. Wadsworth, W. H. Dresher, W. M. Fassell

HIGH temperature-high pressure techniques have long been used to great advantage in the organic chemical industry, the petroleum industry, and the paper industry. Only recently, however, have these methods been used to extract metals from their ores on a commercial scale. The Chemical Construction Corp., together with interested producers of nickel and cobalt, has done much to develop methods of producing nickel and cobalt powders by the use of high temperature-high pressure techniques.' The Howe Sound Co. is currently operating a plant near Salt Lake City in which the new Chemical Construction Corp. technique is applied to cobalt-arsenic sulfide concentrates. In this method the concentrates are leached in a dilute solution of sulfuric acid under air pressure at a high temperature. Near Edmonton, in Alberta, Canada, the Sherritt Gordon Co. is also using the new method in conjunction with its ammoniacal leach process.' National Lead Co., at Fredricktown, Mo., is operating a process similar to that of Howe Sound. In spite of the rapid growth of high temperature-high pressure hydrometallurgical processes, there has been very little work done on fundamental principles involved in this type of process. Conse- quently the literature lacks any significant data in this field. Oxidation of pyrite,3 sphalerite,4 galena," and recently molybdenite" in alkaline solutions has been reported to some extent. This investigation was initiated, therefore, to broaden the application of high temperature-high pressure hydrometallurgy and also to contribute some insight into the basic mechanisms of oxidation in an aqueous medium under these conditions. In addition to the immediate consequences of these processes much valuable information remains to be learned about fundamental reactions when they are carried out at temperatures and pressures slightly higher than those normally encountered. Leaching of molybdenite by alkaline solutions under oxygen pressure has recently been reported in the Russian literature by E. S. Usataya,6 ho found that leaching rate increased with an increase in the pH of the solution—the strongest effects being observed at a pH of 10—and also with an increase in temperature. Usataya also has reported the formation of a protective film on the surface of the molybdenite when the leaching solution was weakly alkaline, neutral, or acidic. He believes that this oxidation process explains the migration of Mo++ in the oxidation zones of molybdenite ore deposits and the subsequent impoverishment of the ores. Equipment: The high temperature-high pressure reaction unit used in this work was especially designed for kinetic studies of this nature, emphasis being placed on accuracy of measurement and precision of control. As details of the unit's construction and operation have been discussed elsewhere,' only basic features and principles of operation will be discussed at this point.

Jan 1, 1957

EXECUTIVE COMMITTEES OF LOCAL SECTIONS New York Meets first Wednesday after first Tuesday of each month. DAVID H. BROWNE, Chairman, JOHN H. JANEWAY, Vice-Chairman. F. E. PIERCE, Secretary, 35 Nassau St., New York, N. Y. P. A. MOSMAN, Treasurer. LEWIS W. FRANCIS. Boston Meets first Monday of each winter month. HENRY L. SMYTH, Chairman. ALFRED C. LANE, Vice-Chairman. HENRY A. WENTWORTH, Secretary-Treasurer, 60 India St., Boston, Mass. ROBERT H. RICHARDS, ALBERT SAUVEUR. Columbia Holds four sessions during year. Annual meeting in September or October. FRANK A. ROSS, Chairman. RUSH J. WHITE, Vice-Chairman. LYNDON K. ARMSTRONG, Secretary-Treasurer, P. O. Drawer 2154, Spokane, Wash. FREDERIC KEFFER, FRANCIS A. THOMSON.

Jan 9, 1915

ALEXANDER BRYDEN Alexander Bryden, clean of engineers in the anthracite mining field, and a member of the pioneer Pittston family, was claimed by death Wednesday, September 26, 1917. The announcement was not a surprise, as he had been in a critical condition for several days. Mr. Bryden had been in failing health for a year past, but his indomitable will kept him at his accustomed duties until quite recently. Two weeks before his death he attended the regular weekly meeting of the Board of Directors of the Miners' Savings Bank in Pittston, Pa. At that time he apparently felt that his end was fast approaching, as he bade farewell to his associates on the bank board. Until a few days before the end, he continued to give attention to business matters at his home, in preparation for the end, which he met unflinchingly. In the passing of Mr. Bryden, northeastern Pennsylvania loses one of the best-known mining engineers, and the community, a man who, throughout his long life, was actively interested in its social welfare and business prosperity. He was for many years prominent in the management of the mining operations of the Pennsylvania Coal Co., and at the time of his death held the position of consulting engineer of the Pennsylvania and Hillside Coal. Companies.

Jan 12, 1917

By J. R. Coulter, F. O. Case, H. G. Satterthwaite, B. Harden

During the two years that the Henderson plant has been in operation, a number of technical improvements have been made by the staff of Basic Magnesium, Inc., the effects of which were realized subsequent to the writing of the paper by Major C. J. P. Ball (p. 28j, this volume). Pellet Mix The process as originally installed required the use of peat moss in the mixture going to the chlorinators. This had to be imported from Canada at a considerable cost, and, besides being expensive, constituted a serious fire hazard in the dry climate of Nevada. Experimentation looking toward its elimination had been going on for some time and was actively pushed as soon as operations commenced. It was found that a pellet mix containing no peat, which would chlorinate at a satisfactory rate, could be cast as a fluid slurry, which would set up into a hard mass having the desired porosity. For carrying out this operation, equipment similar to that used in the manufacture of phosphate fertilizer at Anaconda, Mont., was installed and proved eminently successful. The dry materials, with peat moss and raw concentrates entirely eliminated, are mixed with magnesium chloride solution in two Pratt phosphate mixers and the resultant slurry is poured onto two 48-in. slowly moving rubber conveyor belts. By the time the slurry has reached the head pulley, the material has set su4iciently to be broken and sized as desired. The discharge from the casting belts is fed to the four rotary kilns in the preparation plant, where it is dried and coked to the proper degree for subsequent chlorination. With the peatless feed the capacity of the kilns was found to be about double what it had been before, so that the four rotaries suffice for a 10-unit operation, thus making it unnecessary to operate the extruders and tunnel kilns. In addition to saving the cost of the peat, the new system can be operated with about 85 per cent less labor than was formerly required in the preparation plant. The new pellet production system has also made it feasible to install centralized crushing equipment with suitable conveyors and a loading station. Containers of 8000 lb. capacity were substituted for small cubicles of 1000 lb. capacity which formerly were used for charging the chlorinators, and trailer trucks were substituted for the small trains originally provided for transportation from the preparation plant to the chlorinators. These changes resulted in a considerable saving in handling and transportation charges. Handling Molten Metal The molten magnesium originally was ladled from the electrolytic cells into small gas-fired brick-lined cars standing between the cell rows. From these cars the metal nes ium. into ''cheeses'' weighing from 150 to 300 lb. After solidification the

Jan 1, 1944

By Louis A. Panek

INTRODUCTION AND SUMMARY With respect to the geomechanics aspects, the primary technical objectives in mining by an undercut-cave method are to achieve a controlled, sustained caving of the mineral body and to remove the fragmented ore with a minimum of dilution from surrounding unmineralized rock, while maintaining stability or control of the rock structure around the access openings. The structurally ideal ore body probably does not exist. If the rock mass is so weak as to cave on a short span, it may tend to be too sticky for easy drawing and handling, or create special problems in regard to support of the access openings. If the ore is strong, caving may be difficult and the caved fragments may be of such a large size as to require special equipment for transferring the ore from the caved zone. Given enough time and money to generate data and conduct trials, engineering and ingenuity can devise an appropriate combination of mine layout, sequence of extraction, and mechanical equipment to achieve a technically successful caving extraction operation to meet the foregoing requirements in many types of deposits. The large capital investment involved, however, reduces the freedom to make major changes once the mine development is well under- way, and the penalties for failure to accurately anticipate operating conditions militate against selecting any but the most obvious candidates for mining by an undercut-cave method. The demonstrated capability to extract large, deep, economically marginal deposits by this low-cost, high-volume method of mining provides an incentive to develop a rationale for predicting the cavability and stability characteristics of a deposit prior to mining, so that the undercut-cave method may be extended to a much wider range of mineral deposit characteristics. The ultimate goal is to establish as explicitly as possible the quantitative interrelationships between the measured rock-mass characteristics, the caving span, the size distribution of caved ore fragments, and the sizes and locations of stable access openings. Lacking an understanding of these relation- ships, a designer may readily change some factor in the wrong direction (e.g., excessively reduce the distance between the extraction level and the undercut to increase the convenience of operations for the undercutting crew, increasing the frequency of repairs to the extraction-level support system) or create unnecessary problems elsewhere in the system by introducing a design change that can achieve only minimal improvement in the factor of direct interest (e.g., unnecessarily complicate the ore-transfer system by changing the orientations of the openings, with- out succeeding in the objective of improving the ground support conditions). Although successful predesign is the prime objective, subsequent modifications in mine lay- out and sequence of extraction operations are inevitable. In developing the modified solution, systematic experimentation based on an understanding of the underlying structural relationships, coupled with monitoring measurements of selected diagnostic structural-behavior parameters, can achieve an acceptable solution in a minimum number of steps, which is far superior to the typical operational trial-and-error approach, in view of the cost of implementing each successive change. Since a drill-core sample of ore rock from a successful undercut-cave operation may exhibit a uniaxial crushing strength in excess of 100 MPa, the caving of such a rock mass is now commonly believed to be ascribable to the presence of discontinuities such as joints or fractures throughout the ore body. An essential part of the present investigation was therefore to characterize the natural discontinuities at each of the test sites by measuring their attitudes and spacings. The term "fracture" is used herein in a general sense to include any planar discontinuity without implication as to its suggested mode of origin. Most, but not all, of the fractures are properly termed joints. As a point of departure we may consider the possibility that the rock mass is transected by three families of joints, each family possessing a distinct orientation, such that parallelepipeds of intact rock are delineated by the jointing. Even if cementation is absent between adjoining parallelepipeds, the undercutting of the rock mass will not necessarily initiate sustained caving--owing to the all-around confinement, an arch may tend to stabilize over the undercut unless prevented from doing so by the failures of key blocks of intact rock. Thus, although the jointing can be assumed to weaken the rock mass, creating preferred directions of

Jan 1, 1981