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By H. M. Chance

In recent years an important function of the mining engineer has been the appraisal of the value of mining-properties required by those planning consolidations of a number of' individual operations. In no field of productive industry have the economic results attainable by consolidation been more clearly shown than in the mining-regions of the United States. In many important mining-centers unification of ownership of the most valuable properties has already partially or completely been attained, and, in other districts, projected absorptions and consolidations await only favorable financial conditions for their consummation. The iron-ore districts of the Lake Superior region, the coal-mines of Colorado, of Wyoming and of Utah, the anthracite mines of Pennsylvania, the Connellsville coking-coal mines, and large areas of bituminous coal-land in Pennsylvania, in Ohio, in the Indian Territory and in other portions of the United States, are now largely controlled by a few corporations operating on a large scale, or by a number of corporations working in harmony because of an affiliation and interleaving of interests. The tact of keen financiers and the diplomacy of men of great administrative ability are the chief operative forces in conceiving and accomplishing such consolidations, but in formulating and maturing them, the services of the mining engineer are often required to supply the facts upon which they must be based. For this purpose the expert is expected to furnish full and complete reports covering the quantity of ore, coal or other mineral available, its value per ton, cost of mining and treatment, etc. In reporting upon gold-, silver-, lead-, zinc-, copper- or iron-

Jan 1, 1905

By Stephen G. Dardaganian

Pore volume compresibilitieu measured in the laborutory on core .samples are reported for typical reservoir .sarrrl.rtorres at reservoir pressure. These cornpressibilities ore different for each reseriloir sample and cannot be correlated to porosity. The compressihilities are a1so functions of presslire. In recent years several authors have pointed out the importance of including pore volume compressibility of the reservoir rock in certain reservoir engineering calculations. Hall' showed that in calculating the hydrocarbon volume of an undersaturated reservoir from the production per unit change in reservoir pressure, the neglect of rock compressibility could, in the extreme case of low porosity rock, lead to results in error by a factor of two. Hawkinsl' and Hobson and Mrosovsky showed that both rock and interstitial water compressibilities must be included to achieve satisfactory accuracy in this calculation. Although rock compressibility is an important reservoir property in these calculations, very few data are available. Hall' made compressibility measurements on a few samples in a small pressure range. No other direct pore volume compressibility measurements have been reported. Geertsma' discussed the elasticity of reservoir rock but did not present any new data. In this paper, rock compressibilities suitable for reservoir calculations are reported for a variety of sandstones typical of U. S. reservoirs. The measurements were made through a range of pressures which is also typical of reservoirs. EXPERIMENTAL PROCEDURE Plugs 1 in. in diameter and 2- to 3-in. long were diamond drilled, 14 from cores from seven oil-bearing or potentially oil-bearing sandstone, one from a quarried sandstone, and another from a shallow, oil-free, pure orthoquartzite. The plugs were extracted in toluene, vacuum dried, and then weighed. The plugs were then saturated with kerosene and reweighed. This gave pore volume. They were then again extracted in toluene. vacuum dried, and jacketed in 0.004-in. thick copper foil. A short length of -in high pressure tubing led from the jacket. The jacketed sample was then saturated with kerosene and placed in the hydraulic pressure cell as shown in Fig. 1. Movement of the mercury slug showed the change in pore volume of the sample under a given pressure change. Temperature was about 72" F for all measurements. Before taking any measurements, the external pressure on the sample was raised to 10,000 psig and kept there for several hours. This formed the thin copper jacket around the plug. 7-he pressure was then removed and the sample left at least overnight, during which time it recovered its original pore volume. An external pressure of 12,000 psig was then applied and held constant. The change in pore volume was noted as the internal pressure was raised in 1,000 psig increments to 10,000 psig. The internal pressure was then reduced to zero, again noting pore volume changes,

By S. S. Cole, H. Sigurdson

The melting points of mixtures of titanium dioxide and other titanates have been reported to a limited extent as binary systems and some results have been reported in conjunction with silicon dioxide. The limited data indicated that a low melting zone might exist in a ternary or quarternary system of CaO-MgO-TiO2-Al2O3, since the eutectics reported in the binary systems of MgO.TiO2-TiO2, Al2O3TiO2-TiO2 and CaO.TiO2-TiO2 were of about equivalent composition of TiO2. The importance of a low melting region with a high titanium oxide value in the developing of high titanium slags is fully appreciated by metallurgists. A comprehensive study was undertaken to establish basic information on low melting titanate mixtures. Compounds in the System CaO-MgO-TiO2 -Al2O3 Preparation of the titanates of CaO, MgO and Al2O3 confirmed reported data that the following compounds could be formed under oxidizing conditions in solid state reactions: CaO.TiO2, 2MgO.TiO2, MgO.TiO2, MgO-2TiO2 and Al2O3.TiO2. Under oxidizing conditions it was not possible to react MgO with TiO2 in mol ratios higher than 1:2 without having unreacted TiO2 in the product, nor was it possible to form calcium or aluminum titanates with a higher mol ratio of CaO and A12O3 to TiO2 than 1:1. By fusing mixtures of Al2O3-TiO2 and MgO-2TiO2 together a series of solid solution products were obtained, which showed an MgO-2TiO2 X ray diffraction pattern shifted to smaller inter-planar spacings. Range of Investigation Data in the literature1 indicated that these titanates had melting points which ranged from 1645 to 1860°C. Since the purpose of the study was primarily to provide useful data for smelting titaniferous ores, the work was restricted to the zone of the system which would include the crystalline phases present in the slags. The explored limits were bounded by CaO.-TiO2-MgO.TiO2-TiO2in the base plane. The system was extended to a fourth component Al2O3-TiO2 since many titaniferous ores contain appreciable amounts of A12O3. In the quarternary system CaO-TiO2-MgOTiO2-TiO2-Al2O3TiO2, a tetrahedron was used to represent graphically the components, with one component at each point of the tetrahedron. In the base plane, mixtures were pre- pared to represent fairly uniform changes of composition over the desired range expressed on a mol percent basis. Al2O3-TiO2 was brought into the system in increments of 10 mol pct up to 40 pct. Mixtures were then made up for each of the Al2O3-TiO2 planes. The ranges employed were then represented by five planes cutting the tetrahedron at 10 mol pct Al2O3-TiO2 intervals. Choice of Equipment Although it was necessary for all smelting work to be done in a strongly reducing atmosphere, it was decided that all melting point determinations should be done under oxidizing conditions. This decision was prompted by the fact that TiO2 reduced to lower oxides in a reducing atmosphere with consequent changes in melting points. The use of a micropyrometer and a platinum strip furnace for work of this type is adequately described in the literature.2 With modifications, equipment was selected which provided a fairly rapid method of determining melting points of refractory oxides. The final assembly, which was used, consisted of a platinum strip furnace and a Leeds and Northrup disappearing filament optical pyrometer attached to a special telescope which magnified the sample about 20 diam. The furnace assembly consisted of a platinum strip 0.005 X 0.3 X 2.2 in. mounted on brass posts on a refractory base. This was enclosed in a black steel shell which had a 2-in. opening at the top for purposes of sighting the telescope. The platinum strip was in series with two 0.04 ohm nichrome resistors connected to the secondary of a 2 kw transformer which supplied 17 volts. The current to the primary was varied

Jan 1, 1950

By C. B. E. Douglas

The following analysis was stimulated by a previous article on evaluation of the water problem at Eureka, Nev., which describes a method using formulas especially devised to calculate flow potential of extensive aquifers characterized by relatively even porosity and permeability throughout. The present discussion submits that the method was unsuitable for solving the kind of problem occurring at Eureka, where the amount of water available, rather than the flow potential, may have been the vital factor. IN an interesting article on evaluation of the water problem at Eureka, Nev.,1 W. T. Stuart describes how a difficult water problem, or one phase of it, may be evaluated by means of a small scale test. Test data are plotted by a method rendering, under certain conditions, a straight-line graph that can be projected to show how much the water table will be lowered by pumping at any specified rate for a given time. A formula is then used to determine the size of opening, or extent of workings, necessary to provide sufficient inflow to enable pumping to be maintained at that rate. At first glance this might seem the answer to a miner's prayer, but a word of caution is in order. It may not be the whole answer. Moreover, results obtained by the method described are reliable only for conditions approximating those assumed. Even where conditions do not meet this requirement, however, it may be possible to draw helpful inferences from the results, perhaps enough to facilitate another approach to evaluation of a problem. The two formulas Mr. Stuart used, the Theis formula and the one developed from it by Cooper and Jacob, were given field checks a number of years ago in valley alluvials by the Water Supply Div. of the U. S. Geological Survey and found to be reliable when the aquifer is very large in horizontal extent and sufficiently isotropic for the test well and observation wells to be in material of the same average permeability as the saturated part of the aquifer as a whole." Extensive valley alluvials, sands, and gravels can be evaluated in this way, and there are even cases in which the method could apply to porous limestones, such as flat beds of very large areal extent that have been submerged below the water table after extensive weathering. These are sometimes prolific sources of water for towns and industries. It is necessary for them to have been above the water table for some geologically long period of time in a fairly humid climate before submergence because the necessary high porosity and permeability, and large reservoir capacity, are the result of weathering, that is, of solution by the carbonic acid (H,CO3) in rainwater formed by the absorption of CO, from the air by raindrops, and this dissolving action must cease when all the H2CO3 has been consumed by re- action with the carbonate to form the more soluble bicarbonate. Consequently this weathering process is largely restricted to a zone that does not extend much below the water level, and submergence is necessary after the weathering to provide large reservoir capacity and good hydraulic continuity. On the other hand, water courses tend to form along faults and fractures in limestone, and to become enlarged by solution, well below water level when, as often happens, fresh meteoric water is circulated rapidly through them to considerable depth by hydrostatic pressure, as through an inverted syphon. Although the reservoir capacity of such water courses is relatively small they may extend far enough to tap more prolific sources. Cavities, and sometimes caves of considerable size, are found in limestones where the acid formed by the oxidation of sulphides has attacked them. This action can take place as deep below water level as surface water is carried by syphonic or artesian circulation, because the oxygen it carries in solution will not be consumed until it reacts with some reducing agent, such as a sulphide. Moreover, the formation of acid and solution of limestone in this way is not confined to the immediate vicinity of the sulphide. Oxidation of pyrite, for example, results in formation of acid in several successive stages, each taking place as more oxygen becomes available, as by the accession of fresh water into the circulation at some place beyond the sulphides. When the acid thus formed attacks the limestone, CO, is liberated and the ultimate effect of the complete oxidation of one unit of pyrite will be the removal of six times its volume of limestone as the sulphate and bicarbonate, both of which are relatively soluble. The reaction may be continued or renewed along a water course far from the site of the sulphides, where the small electric potential produced by contact with the limestone helped to start the reaction. Mr. Stuart refers2 0 caves in the old mining area in the block of Eldorado limestone southwest of the Ruby Hill fault at Eureka, Nev., and to the cavities encountered in drillholes in the downthrown block on the other side of the fault. Although he interprets these cavities as evidence that this formation was sufficiently isotropic (evenly porous and permeable) to give reliable results by the method he describes, they may, in fact, be entirely local conditions. There is reason to think they were probably formed

Jan 1, 1956

While coal was mined in New Mexico in 1860, production was on a small scale during the period considered here, and Table 77 contains all the data found. [ ]

Jan 1, 1942

By J. Nutting, J. W. Cahn

WITH the development of thin film techniques for the direct examination of metals in the electron microscope some new problems in quantitative metallography have become apparent. In order to obtain certain contrast effects in the electron optical image when examining thin films from alloys containing dispersed phases, it is essential to use films of a thickness at least two or three times greater than that of the mean particle diameter of the dispersed phase. From micrographs of such specimens it may be desired to obtain the volume fraction of the dispersed phase and the number of particles of the phase in a unit volume of the matrix. Under these conditions the particles sectioned by

Jan 1, 1960

By George J. Simchuk

Selection of an applicable stripping method and the subsequent determination of appropriate sizes of selected equipment are discussed. The case in point is Utah International, Inc.'s Goonyella mine located in central Queensland, Australia. A detailed study of the various types of stripping methods now practiced throughout the world led to the conclusion that large draglines would be the logical choice for Goonyella. The optimal size of dragline was then determined by first designing a series of cut diagrams using an array of machine and pit geometries followed by a cost comparison of the various sizes of machines, in terms of their effective spoil radius. A dragline with a 51-cu-yd bucket was determined to be of optimum size. Further calculations, incorporating total overburden moved from both the highwall and spoil sides, required that four drag lines, two equipped with 40-cy-yd buckets and two with 54-cu-yd buckets, be used to satisfy the stripping requirements at Goonyella.

Jan 1, 1973

[PAGE PREFACE. A. B. PARSONS 3 OFFICERS AND DIRECTORS 7 STANDING AND SPECIAL COMMITTEES 8 PROCEEDINGS OF 1935 MEETINGS 10 NEW YORK, FEBRUARY 10 CHICAGO, OCTOBER 14 SAN FRANCISCO, OCTOBER 14 HOUSTON, OCTOBER 15 ST. LOUIS, OCTOBER 15 ABSTRACTS OF PAPERS PUBLISHED IN 1935 17 TECHNICAL PUBLICATIONS AND CONTRIBUTIONS, 1935 64 INDEX OF 1935 PUBLICATIONS 71]

Jan 1, 1936

By E. J. Duggan

It is a coincidence that the two plants in this country for treating copper ores by ammonia leaching came into existence at almost the same time, that of the Kennecott Copper Corporation at Kennecott, Alaska, in March, 1916, and that of the Calumet & Hecla Mining Co. in July of the same year. The success of both from their very beginning attests the merits of the process under widely different sets of conditions. As would be expected, in the time they have been operating these two pioneers in this field have greatly perfected their methods and improved their technic. The early difficulties inevitably encountered in undertakings of this nature have been largely surmounted and the process may now be considered to be well seasoned and matured. There have been no other commercial plants built in this country and but one elsewhere, the Bwana M'Kubwa in Northern Rhodesia, for the reason that this method of treatment may be employed on comparatively few ores. However, it has certain advantages, some of which are inherent to sand leaching processes generally and some to ammonia leaching in particular. Among these may be enumerated: (1) the avoidance of fine grinding; (2) the small amount of water required; (4) a high-grade product; (5) little likelihood of deleterious gangue constituents and fouling of solutions. It is true that great progress has been made in recent years in the flotation of the ores that come within the province of ammonia leaching. Nevertheless ammonia leaching has its possibilities and these may be overlooked in the present tendency toward the use of flotation for all ores. Probably the chief requisite to its further extension to suitable ores will be a cheap source of steam. Ammonia, the only reagent required, is unlimited in supply, readily transported and the price that is likely to obtain in the future should not make its cost at all onerous. Choice of Method for Kennecott Ores The Kennecott ores are typical limestone replacement deposits. The milling ores have a copper content of approximately 8 per cent in a limestone dolomite gangue. About half the copper occurs as sulfide, mostly chalcocite with some covellite, and the remainder in the form of the two carbonates, malachite and azurite. A comparatively coarse grind followed

Jan 1, 1934

By C. D. Ramsden

DURING the war years of 1941 to 1945, maintenance of mine locomotives and other mine equipment took the form of rebuilding rather than of renewing. Pur¬chase of new parts became increasingly difficult and mine operators found it necessary to maintain equipment by sal¬vaging old parts. Mine-locomotive wheels, although not representing a high cost per pound of material, contain a con¬siderable amount of steel and without rebuilding would have to be scrapped long before even a moderate percentage of the steel in the wheel is actually consumed. Many mine operators have been re¬building locomotive wheels for several years with automatic open arc-welding equipment. In 1927, the Pittsburgh Coal Co. started such a program, using a single Westinghouse automatic welding head. Since that time, many refinements have been added, until in 1944 the setup was operating very efficiently. A desire to increase wheel life and to hold down rising costs led the company in 1944 to investigate the possibilities of the under¬flux arc-welding process for wheel rebuild¬ing, although at that time, little or no work had been done on multiple pass or "build-up" welding with the underflux arc. Work with this process had been confined largely to welding together flat plates in a "down hand" position. Possibilities for savings were intensified by the fact that the company has 300 locomotives in operation, using actively or as spares 1200 sets of wheels. During 1942 and 1943, about 500 sets of wheels were sent each year to the company's maintenance shop at Library, Pa., for rebuilding. Based on having 800 sets in service (many of the company's heavy locomotives have three- sets of trucks), the average life of a set of wheels was approximately 19 months. Anything that could be done to increase wheel life promised sizable annual savings. OPEN ARC WELDING With the open arc process, a set of wheels could be rebuilt for about 40 pct of the cost of new wheels. Underflux arc welding gave promise of greatly increased welding speeds, more homogeneous, dense deposits and the chance for use of a wider range in welding-wire compositions. Against this was ranged the higher cost of materials and the cost of developing the process from the theory to a practical reality. Since 1927, when a single open arc head was placed in operation, fixed welding procedures had been gradually developed, a positioner for handling a set of wheels was built, and a second head was placed in operation in order to weld two wheels simultaneously. The fixture, or posi¬tioner, which was built to position and turn the wheels under the two welding heads, included a motor drive, controlled by a rheostat, which rotated the wheels through a double set of speed reducers and a removable jaw coupling attached to one end of the axle of the wheel truck.

Jan 1, 1946

By R. K. Waring, E. C. Truesdale

The Research Division of The New Jersey Zinc Company (of Pa.) has conducted, over a period of years, numerous tests of the reducibility of various zinc ores and the reactivity of various kinds of coal, using a number of different experimental methods. Most of these tests have involved the reduction of mixtures of ore and coal. In such cases the following simultaneous reactions are believed to be mainly responsible for the production of zinc, as was shown by Bodenstein1 some years ago: ZnO+ CO = Zn + CO2 [1] C + CO2 = 2 CO [2] Each of these reactions is dependent upon the other for the supply of CO and CO2 , respectively, and the over-all rate of formation of zinc vapor will be determined by the slower of these two reactions. An analogous pair of reactions sometimes of importance is: ZnO + H2 = Zn + H2 O [31 C + H2O = CO + Ha [4] While it is possible to obtain an indication of the relative reducibility of various ores, or of a fine versus a coarse ore, by testing them with a given coal, and while coal reactivities may be measured in a corresponding manner, a more direct determination of relative reaction rates is possible by carrying out the desired reactions, such as I and 2, separately and under reproducible conditions. It is one purpose of this paper to present some of the results that have been obtained by the latter method. C. G. Maier2 discusses the relative rates of reactions I and 2 and concludes that while at relatively low temperatures reaction I is faster than reaction 2, this difference becomes less with increasing temperatures, although at smelting temperatures reaction I is still the faster. He also says (p. 48), "At temperatures . . . between IIOO° and I3oo°C., . . . the zinc oxide reduction reaction is intrinsically more rapid than its physical limitations; that is, it is limited by gas diffusion rates." In the course of numerous reactivity tests involving the separate use of reactions and I and 2, we have obtained data that show by direct experiment, we believe for the first time, the correctness of Maier's statements concerning these reactions. We also include data on the relative rate of reaction 3, involving reduction by hydrogen. Finally, some interesting results obtained by carrying out the reduction of a zinc ore-coal briquet in streaming atmospheres of nitrogen, carbon monoxide and hydrogen, respectively, are presented. *

Jan 1, 1943

By E. C. Truesdale, R. K. Waring

The Research Division of The New Jersey Zinc Company (of Pa.) has conducted, over a period of years, numerous tests of the reducibility of various zinc ores and the reactivity of various kinds of coal, using a number of different experimental methods. Most of these tests have involved the reduction of mixtures of ore and coal. In such cases the following simultaneous reactions are believed to be mainly responsible for the production of zinc, as was shown by Bodenstein1 some years ago: ZnO+ CO = Zn + CO2 [1] C + CO2 = 2 CO [2] Each of these reactions is dependent upon the other for the supply of CO and CO2 , respectively, and the over-all rate of formation of zinc vapor will be determined by the slower of these two reactions. An analogous pair of reactions sometimes of importance is: ZnO + H2 = Zn + H2 O [31 C + H2O = CO + Ha [4] While it is possible to obtain an indication of the relative reducibility of various ores, or of a fine versus a coarse ore, by testing them with a given coal, and while coal reactivities may be measured in a corresponding manner, a more direct determination of relative reaction rates is possible by carrying out the desired reactions, such as I and 2, separately and under reproducible conditions. It is one purpose of this paper to present some of the results that have been obtained by the latter method. C. G. Maier2 discusses the relative rates of reactions I and 2 and concludes that while at relatively low temperatures reaction I is faster than reaction 2, this difference becomes less with increasing temperatures, although at smelting temperatures reaction I is still the faster. He also says (p. 48), "At temperatures . . . between IIOO° and I3oo°C., . . . the zinc oxide reduction reaction is intrinsically more rapid than its physical limitations; that is, it is limited by gas diffusion rates." In the course of numerous reactivity tests involving the separate use of reactions and I and 2, we have obtained data that show by direct experiment, we believe for the first time, the correctness of Maier's statements concerning these reactions. We also include data on the relative rate of reaction 3, involving reduction by hydrogen. Finally, some interesting results obtained by carrying out the reduction of a zinc ore-coal briquet in streaming atmospheres of nitrogen, carbon monoxide and hydrogen, respectively, are presented. *

Jan 1, 1943

By W. G. Woolf, E. R. Crutcher

The electrolytic zinc plant of the Sullivan Mining Co. is in Government Gulch, at Silver King, Shoshone County, Idaho, in the Coeur d'Alene mining district, about one mile from the Bunker Hill smelter and about three miles from Kellogg. The Sullivan Mining Co., which owns both the zinc plant and the Star mine, is itself owned jointly by the Bunker Hill & Sullivan Mining & Concentrating Co, and the Hecla Mining Company. The Tainton high-density, high-acid process is used at this plant, therefore some parts of its flowsheet differ markedly from those of electrolytic plants using the low-density process. While it is a custom plant and has treated a variety of zinc concentrates originating from various sources in a number of states, it has obtained its main supply of zinc concentrates chiefly from the zinc-bearing ores of the Coeur d'A1ene mining district, of which the principal important sources have been the Bunker Hill and Star mines. Thus the main feed has been zinc concentrates obtained from complex lead-zinc-silver ores. For the year of 1935 the concentrates had the following composition: water, 9.5 per cent; Zn, 50.6; Pb, 3.3; Fe, 9.4; S, 31.0; Cu, 0.2; insoluble, 2.9. Flowsheet The concentrates are delivered to the plant in railroad cars, which discharge into storage bins directly under the track. Bottom-dump cars are dropped directly into the bins while a slip scraper, pulled by an air tugger hoist, is used to empty box cars. There are five storage bins for concentrates, each of approximately 300 tons capacity. At the bottom of each bin there are individual belt feeders that discharge onto a common conveyor belt, which in turn discharges to another belt that carries the concentrate onto a weightometer belt for weighing. Immediately after the weighing, the crushed oversize return from the roasters is added to the load on the belt and then the ground-up sweepings, lumps and roaster cleanings. An electromagnet at the end of this conveyor removes tramp iron from the feed. The belt discharges to a set of 30 by 12-in. rolls, which break up lumps or frozen chunks of concentrate, etc. Beneath the rolls there is a screw conveyor, which conveys the roll discharge to a

Jan 1, 1937

By T. Reed Woodbridge

The object of this paper is to bring to notice a method in use at the Cedar Point laboratory of Witherbees, Sherman & Co., for determining the phosphorus in the ore mined at Mineville, N. Y. This ore is magnetite, with a gangue containing crystals of apatite. In 1886, Mr. H. F. Dawes, then head chemist at the works, suggested that as all the phosphorus seemed to be in the apatite, it could be obtained in solution by dilute nitric acid. After some experimenting the following method was adopted, and has proved invaluable for rapidity and ease; half a dozen determinations requiring very little more time or trouble than one. From + to 10 grammes of ore are weighed into a 150 c.c. beaker; 50 C.C. hot water and 10 c.c. strong nitric acid are added; the beaker is covered and the contents are boiled for 1/2 hour (10 minutes sometimes proves sufficient; but 4 hour is safer). The bumping which ensues if more than & gramme is used need cause no trouble. The hot solution is filtered into a 250 C.C. beaker. The filtration and washing (with hot water) proceed rapidly, as the magnetite itself is but slightly acted upon. After the addition of 15 C.C. ammonia and 15 c.c. nitric acid, the solution is cooled to 45' C., and from 15 to 40 C.C. of molybdic acid solution is added and stirred vigorously. In 5 minutes the precipitate will be down, and the determination can be finished by any good method. At Cedar Point the method more commonly used is weighing as yellow salt. Good washed filter-paper is used, No. 0 Swedish being preferred; and dozen 3 1/2-inch filters are numbered and dried and weighed at 110°C. These are always kept in stock. The yellow salt is filtered on one of these, washed with dilute nitric acid (2 per cent. by volume); dried for 1 hour at 110' C., and then weighed, 1.63 per cent. of weight being calculated as phosphorus. The filters vary in the two weighings from

Jan 1, 1889

By W. H. Meyer

During the seventeen years Fairchild has been making aerial surveys and aerial photographic equipment many changes and improvements have been made in the equipment and in the technique of using it. Aerial photographic operations have been developed from the crude methods and apparatus of wartime to a science upon which exact engineering calculations can be based. In the early days there was no airplane specially designed to meet aerial photographic requirements. One type of camera was used in all types of operations. Projects were made to conform to the products we could make and deliver. Obviously this entailed definite limitations. We now have airplanes designed specially for aerial photographic operations in accordance with the conditions encountered in actual service experience. We have cameras for every photographic purpose and our laboratory technique and equipment have been developed to the point where it is possible to make contour maps from vertical aerial photographs with contour intervals as close as five feet. Therefore, the selection of equipment now can be made in accordance with the time at our disposal, the justifiable cost, the scale desired, and the general requirements of the client. This does not mean, however, that arbitrary deliveries can be set for aerial photographic products. The factors in each project vary and each job must be considered in accordance with the factors involved. This accounts for the fact that products of aerial surveys may readily vary from $1.25 to $200 a square mile. Sherman M. Fairchild became interested in the design of a radically different aerial camera during the war. At its close he was about ready to demonstrate his design to the War Department. By 1920, he was able to bring out his first completed model. It had a number of radical features, one of which was the first dependable high-speed between-the-lens shutter, which makes for accuracy in aerial photographic work. Another feature was the use of metal throughout, eliminating the wooden parts that had been employed in wartime cameras. Another feature was the interchangeability of parts.

Jan 1, 1937

By R. E. J. Putman

Introduction Self-tuning regulators are a class of controllers built around the concepts of sampled data theory, and are especially suitable for implementation in a digital microprocessor or minicomputer control system. The algorithms used are distinctly different from those based upon analog control techniques and include automatic adjustment features which permit them to adapt to unmeasured changes (in, for example, ore grindability) or process nonlinearities (as exhibited by cyclone classifiers). They are also designed to handle transport lags in a simple manner, for example, anticipating the effects of a variable recycle rate on particle size distribution. The advent of distributed control systems will encourage the exploration of these techniques in the minerals processing industries, and particularly for improving the control of comminution circuits, which are especially energy intensive. This paper outlines some current work in progress on this topic. Background Those engaged in the practice of process control have long had two philosophical problems for which solutions have been actively sought - what is optimal control and how can we attain it in some heuristic fashion automatically. The basic analog controller can be represented by the differential equation: [V = Ko + K1 f y • dt + K2 dt(1)] where: v and y are the instantaneous values of output and error respectively and K0, K1, and K2 are tuning constants. The empirical rules for establishing the values of K0, K1, and K2, which provide stability, have evolved from the initial suggestions of Ziegler and Nichols(9) in 1942 and have served their purpose well. However, later analyses demonstrated that there are in fact many combinations of the values of K0, K1, and K2 which, besides being stable, behave better or worse in relation to other process control criteria; some examples are duration of error, amplitude of a damped response, magnitude or rate of movement of control valve or adjustment of speed regulator, etc. The definition of optimality had broadened. In 1964, Kalman(4) published his classic paper which tried to resolve these problems. He defined a

Jan 1, 1982

By K. Stout, E. Van Eeckhout, L. Friel

Previous research at the Montana College of Mineral Science and Technology indicated that it is practical and economically feasible to place stiff back- fill in a cut-and-fill stope. This backfill would reduce wall closure and high stress buildup in pi1llars, thus alleviating rockburst potential. Three ;topes from the Sunshine Mine, of the Coeur d’Alene Mining District, were chosen for a test verification of this concept. Although the first stope was lever mined beyond the initial cut (for economic reasons), some vibrating wire stressmeter results were obtained and found to compare favorably with numerical predictions. Some other instrumentation devices were built and tested, but never installed .

Jan 1, 1984

By P. G. Shewmon, F. N. Rhines

THE determination of the self-diffusion coefficient of magnesium has been made possible recently by discovery1-1 of a radioactive isotope, Mg28 having a half-life of 21.3 hr,1 and subject to manufacture in useful quantity. In the present research this material was condensed from the vapor phase upon a surface of high purity magnesium. The progress of diffusion of the tracer atoms into polycrystalline magnesium was followed by machining layers and measuring the change in the intensity of radiation as a function of the distance of each layer from the surface. The self-diffusion coefficient was found to be 2.1 X 10-8 sq cm per sec at 627°C, 3.6 X 10-9 sq cm per sec at 551°C, and 4.4 X 10-10sq cm per sec at 468°C; the activation energy is about 32,000 cal per mol. Experimental Procedure Since there was no other published measurement of a diffusion velocity in any magnesium-base material, is was necessary to employ a number of new experimental techniques. The short half-life of Mg28 made it necessary to complete the entire experimental procedure within three or four days. This meant that the work had to be done where a cyclotron was readily accessible and that all operations, prior to the diffusion heat treatment, had to be so designed as to minimize their time requirements. Unusual problems were imposed also by the chemical reactivity of magnesium, its high vapor pressure, and the fact that no satisfactory method for elec-trodepositing magnesium on magnesium is presently available. Finally, the machining and handling of the easily air-borne radioactive-magnesium chips involved certain health hazards, resulting in the need for further experimental restrictions. Preparation of Mg28 The Mg28 was produced in the Carnegie Institute of Technology syncrocyclotron by the neutron spallation of chlorine.5 his involved bombarding a 2 gram crystal of high purity NaCl with a beam of 350 mev protons for a period of 2 hr, after which the crystal was dissolved in warm water and the Mg28 was concentrated and purified by chemical means (see Appendix). About 50 microcuries of Mg28 thus were obtained in the form of magnesium oxinate (8 hydroxyquin-olatc?), which was ignited in air to produce MgO. This in turn was reduced to magnesium metal vapor, by the method of Russell, Taylor, and Cooper," in the vacuum apparatus shown schematically in Fig. 1. Here the essential part is a tantalum ribbon, slightly dished to receive the MgO. The ribbon, pre- viously outgassed at high temperature, is heated to about 1700°C by passing an electric current through it, whereupon tantalum oxide is formed, magnesium vapor is released almost instantaneously, and condensed partly upon the diffusion sample. Diffusion-Sample Preparation: Hot-extruded magnesium rod, 21/32 in. round was used in making the diffusion specimens. The magnesium analyzed as follows: 0.004 pct Al, 0.027 pct Fe, 0.040 pct Mn, 0.0004 pct Cu, 0.0002 pct Ni, and less than 0.01 pct Ca, 0.0004 pct Pb, 0.0011 pct Si, 0.001 pct Sn, and 0.001 pct Zn. A brief study of the crystal texture of this material revealed a sharp fiber texture with the (001) plane roughly parallel to the extrusion axis. Cylindrical samples 1/2 in. long by 5/8 in. were machined from this rod, the end faces dressed on 3/0 emery, and lightly etched with 20 pct HC1 in water. These samples then were annealed for at least twice the intended time of diffusion, at the intended diffusion temperature, in order to stabilize the grain structure at about 1 mm average diameter. The annealing treatments were conducted in argon in the same apparatus and in the same manner as the subsequent diffusion treatments, which will be described presently. Thus, a strain-free plane surface was produced, but there remained a layer of MgO which had largely to be removed before the layer of Mg28 was deposited. Most of this layer was taken off by two light passes over 3/0 emery paper. The balance of the oxide and a thin layer of metal were then removed by etching 5 to 10 min in 4 pct nital (4 pct HNO3 and 96 pct ethyl alcohol) made with absolute alcohol. There followed immediately three quick rinses in: 1-49 1/2 pct methanol, 49 1/2 pct acetone, and 1 pct formic acid, 2-50 pct methanol and 50 pct acetone, and 3-pure benzene. This procedure is essentially that of Sturkey.7 The resulting surface, which was of almost elec-tropolished brightness, remained plane and was free of cold work. It could be kept clean by storing under benzene, or in a desiccator; short exposure

Jan 1, 1955

By Caner Zanbak, Erdogan Yüzer, Mahir Vardar, Kemal A. Erguvanli

The Hasancelebi iron deposit in Eastern Turkey is a very low grade (average 20% oxide) magnetite body consisting of 720 million longtons of proven reserve at a 10% cut-off limit. The ore-bearing scapolitic oclcs cover an area of approximately 4300m x 300m, and the thickness of the outcropping deposit exceeds 300m. Prefeasibility project studies revealed that a one degree decrease in a 45° final slope angle of the proposed pit would require excavation of approximately 10 million cubic meters of excess waste material, for a projected pit bottom 300 meters below the valley floor. There- fore, determination of the final slope angles of the open pit was considered to be the most important task in the feasibility analysis of this project. In this investigation, the main emphasis was given to collection, compilation and evaluation of the engineering geological, hydrogeological and rock mechanics data of the highly fractured and weathered rocks. Due to the presence of intense fracturing, the rock wedge analysis method was used for projected 30m-high benches and the circular failure analysis method was used for the stability of the projected final pit slopes. The proposed bench slope angles ranged between 60° and 80°, whereas the projected final slope angles varied between 30° and 48°, depending on the slope heights and groundwater levels.

Jan 1, 1983

By Bernard N. Moore

At present the demands of the United States for strorltium are met by imports from Germany, England and Canada, which vary considerably in proportions of ore and finished salts, in tonnage and in value. Statistics for the period 1926-1932 are given in Table 1. Domestic deposits arc known in many states, but the larger and purer ones are in the West, particularly in the desert region of southeastern California and western Arizona. These deposits are large enough to command interest, particularly in consideration of future cheap power at the Boulder Dam. The deposits described in this paper were examined during an investigation of the mineral resources of the Boulder Dam region by the United States Geological Survey for the United States Bureau of Reclamation in the spring of 1934. Mr. D. F. Hewett, who had general charge of the investigation, Mr. W. T. Schaller and Mr. M. I. Goldman contributed many helpful suggestions. Mr. N. H. Harshman assisted efficiently during the field work. Celestite Deposits The celestite deposits described in this paper are of sedimentary origin and Tertiary age. They are widely scattered, as shown on Fig. 1,

Jan 1, 1935