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By J. D. Price, W. M. Bertholf

Making maximum possible use of available equipment and material, CF&I placed a high-efficiency, high-capacity washery unit in the existing buildings to gain simplified operation, reduced manpower requirements, raised efficiency, and boosted production. THE central washing plant of the Colorado Fuel & Iron Corp. was first operated in 1918 to furnish coal for two 60-oven batteries of Koppers design. Prior to that time the coke for the blast furnaces had been secured from a number of beehive operations, some with their own washeries. The original washery had four Elmore jigs, which did not do as well as had been expected. In 1923 the jigs were taken out and replaced by Deister Plat-O tables, with various changes in auxiliary equipment. In 1936 the throughput of the plant was considerably increased by the installation of vibrating filters for recovery and drying of fine coal, thereby reducing a large recirculating load of fine coal which passed through the screens of the Carpenter driers. In the late 1940's it became evident that something would have to be done to keep the washery abreast of progress at Pueblo and other washeries. Coal requirements were nearly two and one-half times the original requirements of the 120 small ovens, as additional ovens had been built in 1930, 1938, and 1945, whereas no additional coal washing capacity had been added since the 1936 modifications. Much of the equipment, moreover, had seen better days. One answer to the problem was to build a new and completely up-to-the-minute plant, with sep- arate sections for coarse, medium fine, fine, and very fine coal. On the other hand, it might be financially more profitable to place a high-efficiency, high-capacity unit in the existing buildings. The washing problem at Pueblo is not unique, but it is approached differently. There are no established specifications for coke ash and sulphur. The plant is operated on the principle that if it is more economical to wash it out than to slag it out, wash it out; otherwise, send it on to the blast furnaces. With coals at the Pueblo plant costing what they do, the benefits of washing at a low gravity are questionable. Hence the primary problem is to eliminate the high-ash components and retain all the low and medium-ash components. Sulphur is no problem at all, averaging 0.6 pct in the coke. Since a good proportion of the free impurity in the coals is as large as the coal or bone or larger, it appeared that it would be safe to go to jig washing and discontinue crushing the coal before washing, as had been the practice with the tables. This would boost the throughput. It might also decrease manpower requirements and washing costs considerably. It was assumed at the start and has been demonstrated in the first year of operation that a jig will do as well on the fines as the tables had been doing. It would not be necessary to discount any gains due to increased efficiency in washing the coarse material because of incapacity to wash the fines. By washing at —2+0 mesh it would be possible to dewater, by screening, a considerable tonnage satisfactorily, for example, at —2 + % or —2 + l/4, and then crush it in a hammer mill. The —?+20 could be dewatered in centrifugal driers and the remain-

Jan 1, 1955

By O. E. Palasvirta

Close examination of most so-called n.eu processes in mineral dressing reveals that they were conceived and developed a long time ago. High-intensity magnetic separation is no exception. Although its application in iron ore beneficiation seems new, in Germany the process has been commercially successful for almost two decades. High-intensity separators have long been standard equipment in the glass sand industry; they have also been used in concentrating some of the more unusual para-magnetic minerals. The basic features of an induced-rotor, high-intensity magnetic separator were described by Desire Korda in Paris in 1905,' and many patents have since been issued for improvements on the original design. Fig. 1 shows the basic design of an induced-rotor separator. Several U. S. manufacturers have been supplying induced-rotor separators for many years, but these are all devices of relatively low capacity, not suitable for treating such low-cost commodities as iron ore. On the other hand, two German manufacturers are offering induced-rotor separators of high capacity, both proven in the field. The present investigation was prompted by the apparent success of the German installations and by the discovery of iron ores on this continent which may be amenable to high-intensity separation. The objective was to gain better insight into the underlying principles of separation in a high-intensity field and to use this knowledge in selecting or designing an optimum separator. THEORETICAL CONSIDERATIONS he well known low-intensity magnetic separators are. without exception, of the the magnet type with more or less similar field shapes. It is therefore permissible to speak of their comparative strengths in terms of field intensity alone. However, proper understanding of the high-intensity magnetic separation process requires somewhat deeper analysis of the relationship between the magnetic field and the attractive force. Contrary to a widely held belief, a magnetic field of parallel lines of flux exerts no attractive force on a magnetically susceptible particle, but merely aligns its poles parallel to the lines of flux. On the other hand, a non-uniform field attracts a magnetically susceptible particle in the direction of the converging lines of flux. Thus the force of magnetic attraction varies as the product of the field intensity and the field gradient: dH ds As the induced-rotor magnetic separator is primarily intended for separating feebly magnetic materials, it is important that maximum attraction be attained. For the designer, then, the relationship between the magnetic field and the attractive force has the following implications: 1) Insofar as possible, the separation gap must be saturated with flux. 2) The rotor surface must be shaped to insure maximum convergence of lines of flux toward the rotor. 3) The primary pole face (opposite the rotor, on the other side of the separation gap) must be shaped so that the lines of flux do not diverge from it at any location in the effective separation zone. The actual attractive force of the rotor, and its critical speed, can be calculated as shown by Runo-linna, provided that the true flux density, the field gradient, and the magnetic susceptibility of the mineral can be measured with sufficient accuracy. The irregular shape of the rotor surface makes flux measurements difficult, but it can be done if a suitable gaussmeter is available (see Fig. 2). Calculations by the author, and comparisons with the centrifugal force exerted by the rotor, indicate that the force pinning a hematite particle to the rotor is of the same order of magnitude as the force holding a

Jan 1, 1960

By Charles Legrand

The power plant of the Burro Mountain Copper Co. is located near Tyrone, N. M., at 5,950 ft, elevation. It is interesting because it uses the largest stationary Diesel engines in the United States. The general layout of the plant is shown in Figs. 1 and 2. The building is of steel-frame construction with hollow-tile walls and corrugated-iron roof; the floor is concrete and all windows have steel frames, making a fireproof building. The building is made to accommodate one more electric generator unit, which is being installed; also two 4,000-cu. ft. air compressors direct-connected to three-cylinder engines with the same size cylinders as the electric units. The power plant consists of two 815-kva., 60-cycle, three-phase, 6,600-volt, 180-r.p.m. generators, direct-connected to five-cylinder, two-cycle Diesel engines; the necessary circulating pumps for cooling water, cooling towers and oil storage tanks. A portion of the power is used in the same building as the generating units to run one of two 200-kw. rotary converters delivering 260-volt direct current, and a 2,500-cu. ft. air compressor, delivering air at 95 1b.-pressure, direct-connected to a 400-hp., 6,600-volt, three-phase synchronous motor. The remainder of the power is used to run a 1,500-ton concentrating mill 3 1/2 miles from the power house, various hoists, pumps and motors about the mines and the lighting and water-works of the town of Tyrone. The direct current is used for electric traction in the mine, also to run hoists and motors that were installed before the new power plant was built. Due to the large hoists that take starting peaks of 250 kw. and the 10-ton electric locomotive that takes at times 150 kw., the load is extremely variable, as shown in Fig. 3. The two five-cylinder vertical Diesel engines, which are rated at 1,250 b.hp. at sea level, have cylinders of 525-mm. diameter (20.6 in.) and

Jan 1, 1917

By Clarence Zener

THEORETICAL considerations by one of the authors have led1 to the prediction that the dynamic internal friction of annealed metals has a broad maximum at a certain grain size. This prediction they have veri-fied2 for alpha brass. In the present paper the theory is extended to include the precise manner in which the internal friction varies with grain size on either side of the maximum. New experimental data are pre¬sented to test this extended theory. The physical basis for the type of internal friction that is being investi-gated lies in the interplay between stress and temperature. Just as local fluctuations in temperature give rise to local fluctuations in stress, so like-wise fluctuations in stress give rise to fluctuations in temperature. But temperature gradients give rise to necessarily irreversible thermal cur-rents, and hence to internal friction. The theory for this type of internal friction has been developed quantitatively,3 and has been experimentally verified,4,5 for the particularly simple case of the transverse thermal cur-rents accompanying transverse vibrations. Of more fundamental impor-tance are the intercrystalline thermal currents accompanying all types of vibration. The variation in stress across grain boundaries arises from the elastic anisotropy and at least partial random orientation of the individual crystallites. An exact quantitative treatment of the internal friction due to these thermal currents is of course out of the question. However, much information can be obtained from general arguments.1,6 Thus when the frequency of vibration f is low, or when the grain size d is small, adja-cent grains remain nearly in thermal equilibrium, and vibration proceeds isothermally with little internal friction. Again, when the frequency is high, or the grain size large, the heat flow per cycle is small, the vibration proceeds adiabatically with little internal friction. The internal friction will become a maximum in the transition region. The degree of adiabatic-

Jan 1, 1940

R. M. Raymond, New York, N. Y.—The filling and drawing down of the overhead material was done at considerable depth, which is not the usual method in which it starts at the surface.. R. J. Mechin.— It is a little more than 1000 ft. at El Bordo. The vein above had been caved practically to the surface and some of the ore had been excavated, using underhand and overhand square-set methods. This was far enough away from the shaft to prevent any bad effects. When the sill floor was taken out, to insure the old cave material following down well, we put chutes in the square sets for several weeks until we were sure that the material from above was moving down; on each level above, several crosscuts were driven into the old fill by which we could see the material coming down and thus be sure it was following the first. We had to dynamite it very little because it was dry and dusty. The only trouble was that timber starting from the old square sets would get across the chute and hang up. R.. M. Raymond.—HOW many floors did you operate at once? R.. J. Mechin.—We would operate one floor through the whole length and prepare for the second. F. W. SperR, Houghton, Mich.—Could you have worked it in steps? R.. J. Mechin.—Yes; but working it all on one floor was much better for ventilation. The passage between steps would tend to become heavy and blocked and, as the passages from the mining floor down are usually filled with ore, there would be no way for the air to pass up. By means of doors and fans, air was driven to the extreme end of the orebody through the stope and up the other end; but it could have been worked in steps. F. W. SpeRR.—D the following-up material ever become so heavy that you could see weight on the caps before you had to take out the posts?

Jan 1, 1927

By Albert E. Reed

THE development of the modern high-speed vibrating screen, together with the increasing availability of long-lasting stainless-steel screen cloth for relatively fine-mesh separations, means that more screen circuits will be used in the future. It seems pertinent therefore to examine a few characteristics of screen-circuit products, and especially the points of difference as compared with products of classifier circuits. The most noticeable change that takes place when substituting screening for classification in a grinding circuit is a substantial drop in the circulating load, owing to the greater ease with which the "fines" leave the circuit. Where circulating loads of 500 to 600 per cent are considered fairly normal in classifier grinding circuits, the grinding circuit equipped with screens for equivalent work will have a circulating load of only 20 to 75 per cent. As a result, the products of screen circuits will show a minimum of overgrinding or sliming as compared with a classifier circuit. The desirability of this characteristic will naturally vary according to the metallurgical requirement in each case. Because screen separations make a sharper division according to particle size, the products of screen circuits can be expected to show greater crowding of particles near the point of separation. These changes in product characteristics are so important that other factors, both in the grinding and concentration processes must be adjusted if the operator is to take full advantage of the increased tonnages and higher efficiencies available to him through the use of the screen circuit. One factor that too often is overlooked when changing to screen circuits is the importance of adjusting the ball-mill or rod-mill charge, to compensate for: (I) increased tonnage of new feed, (2) decreased circulating load, or (3) more granular character of feed. These same factors should also be taken into consideration for their effect on the various types of concentration or flotation equipment that may follow the screen circuit. Over a period of years, this equipment has generally been designed and adjusted to the product of the classifier, therefore it seems obvious that some research on the adjustment and adaptation of concentrating and flotation equipment to screen-circuit products will be most profitable to the operators. To illustrate the points mentioned above, the results from three recent installations follow, with data on the comparative characteristics of screen and classifier circuits. A range of separation from 10 to 48 mesh has been chosen, using two general types of screening equipment and two types of classifiers. The first of the installations of which the results are to be examined is that of an iron-ore installation where eastern magnetite was being concentrated prior to sintering. Fig. I shows the original grinding circuit, using classification, while Fig. 2 shows the same circuit equipped with a

Jan 1, 1945

By A. Eilers

Although the process to which I refer in this paper has been in successful operation for nearly five years, during which time it has been introduced, superseding all other processes having in view the same end, in almost every lead refinery in the country, no description of it, giving economical data, has ever appeared in print, nor has the process been known in its details to any considerable number of engineers beside those conducting it at their works. A discussion of this subject at the present time becomes especially interesting in view of a late publication in the Prussian ministerial organ for Mining and Metallurgy, by Dr. Wedding, in which the method of treating the zinc-silver-lead alloy by distillation, as finally settled upon at Tarnowitz, is very minutely and fully described. It is the more interesting, because in that article reference is made to the American method in an unfavorable light, while, at the same time, it is clear from the article itself, that the most important details of the process, besides the one that it is conducted in black-lead retorts, mere not at all understood. In publishing herewith the details of our American method, one of my objects is to invite comparisons between our simple and direct process and the elaborate one at Tarnowitz. In extracting the silver from work-lead by means of zinc, it is always the intention, in this country, to obtain a zinc-crust so rich in silver that the rich lead resulting from subsequent distillation

By F. L. Clerc

The lead and zinc region of SW. Missouri is interesting, not only by reason of the value of its output, which ranges in the neighborhood of ten million dollars a year, but even more because of the facilities which it offers for the study of certain forms of ore-deposits, of' which the loci and genesis are somewhat obscure. Although the mineral species found in the region are few and of common occurrence, crystallization has taken place on a generous scale, and unusual forms abound in many combinations, presenting examples of pseudomorphism, para-genesis, metasomatic replacement and the action of mineral-depositing and mineral-dissolving waters. For the mineralogist, lithologist, and specialist in certain lines of metamorphism, it has peculiar attractions, while the systematic geologist will perhaps find its chief claim to attention in the simplified problem it presents, in the study of ore-deposits. In addition to the phenomena above enumerated, it presents instances of well-developed comb-structure in the ores, fissures and faults, and slickensides—features usually associated with extensive movements in the earth's crust, and believed by some authorities to prove the deep-seated origin of the ores. In connection with these occurrences, the absence of all igneous rocks from this region is noteworthy and significant. For that reason, the simplified case presented by the Joplin region, from which these rocks have been eliminated as a factor, is specially valuable. Having had unusual opportunities to watch the development of this district, I have been deeply interested in the bearing of the revelations of the Joplin mines upon questions of universal importance to the science of ore-deposits; and it was to my great regret that the project of a meeting of the American Institute of Mining Engineers at Joplin, proposed in 1892, was

Jan 1, 1908

By M. K. Yen, W. R. Hibbard

Previous studies of plastic deformation of metals have emphasized the important role of bending and constraints during strain under relatively pure stresses.1"5 Some new phenomena such as early conjugate slip6 and polygonization7 are intimately concerned with the relief of bending stresses, the former by slip and the latter by a process analogous to re-crystallization. However, few analyses of the bending deformation of single crystals were found in the literature, and those8,9 are sufficiently previous to the modern interpretations of slip to invite further investigation in this area. The transverse bending mode of testing, using the two-point loading method, was selected because the center portion of the specimen may be considered as under a theoretically pure bending stress with a uniform bending moment. Considerable attention was given to both the lattice distortion and flow characteristics of the crystals during deformation. Experimental Proeedare A specimen under the action of equal and opposite couples at both ends is said to undergo pure bending. The shear and moment diagrams of a specimen under the two-point loading method of transverse bending are shown in Fig 1. The moment over the middle-third of the span is equal to PL/6, and the vertical shear is zero. Fig 2 (A) illustrates the stress distribution in the elastic range on a plane across the middle of the specimen. It can be seen that the normal stress has a maximum value at the extreme surfaces on the tension and compression sides, and decreases proportionately to zero as it approaches the neutral plane. However, when the plastic range is reached, the stress distribution is no longer a straight line relation, but changes in a manner which can be best illustrated as shown in Fig 2(B). The following calculation and stress analysis are mainly in accordance with the works reported by Timoshenkol0 and Kochendörfer.9 From the simple beam formula, the maximum normal stress will occur at the surface of the specimen. For a round specimen with a radius r and for a rectangular specimen with a height 2h and width 2b, the maximum normal stress, Sp, will be given by the following formulas: For round specimen: 2PL Sp = 3pr3 [1] For rectangular specimen: Sn = 8bh2 [2] Since it is reasonable to consider that slip takes place under the same conditions as in uniaxial loading, the resolved shear stress S?, along the operative slip direction will be as follows: For round specimen: 2PL sin x cos ? Sp = 3pr3 sin x cos ? [3] For rectangular specimen: PL S8 = 8bh2 sin x cos ? [4] where X is the angle between the specimen axis and the slip direction and X, the angle between the specimen axis and the major axis of the glide ellipse. It was also reported by Kochenclorfer9 that the critical bending moment, that is, the bending moment exerted on the specimen to initiate slip, should be higher than the value given by the equations stated above. Based

Jan 1, 1950

By L. Wahl, M. Allard, B. Trentini

REMOVAL of sulfur in steelmaking is presently a problem of utmost importance as requirements on final sulfur contents in finished steels become increasingly strict. This is in spite of often increased sulfur contents in the fuels used during the different stages of steel processing. According to local conditions or current economic factors, sulfur can be removed during one or more of the processing operations; for instance, in a blast furnace operated under high temperatures and with a high slag basicity ratio and also by means of high inputs of manganese into the burden. These methods of proceeding are, however, expensive and in contradiction with the acid practice which is gradually being developed owing to the fact that it allows coke and lime savings and usually permits an increase of productivity. In the melting-shop, desulfurization can proceed in the open-hearth or electric furnace with one or several very basic slags, or with special prefused slags (Perrin process) mixed with steels poured from the open-hearth, electric furnace, or converter. Liquid pig iron has also been treated separately before being used in the melting shop, in order to reduce high sulfur contents. This external desul-furization has been developed in Europe for the basic Bessemer Thomas process, which is again in favor owing to the great flexibility of such installations and to the fact that the steel produced is not polluted by impurities coming from the scrap. Since the basic Bessemer process, as usually practiced, has a low desulfurizing power, external desulfurization is generally carried out in ladles by means of alkalies such as soda ash. This method, which has the advantage of being simple and costing little, has only a medium desulfurizing power, which implies the necessity for repeating the operation several times when a substantial amount of sulfur has to be removed or when very low sulfur contents are desired. Thermal losses are high and there is a considerable decrease in output. Moreover, the resulting fluid and corrosive slag, which is extremely difficult to eliminate, attacks the refractory linings of the mixers and steel furnaces. Therefore, it seems necessary to devise a simple method for external desulfurization which would not only remove substantial quantities of sulfur from the hot metal coming from an acid blast furnace, but which would also bring the sulfur contents down to the low level required in the making of high quality steel and iron. Principle of the Process In this paper a new method will be described for desulfurizing hot metal. It has been perfected by

Jan 1, 1958

By Granville S. Borden

The fruits of industry are divided between capital, labor, and governments. Capital takes its redemption and remuneration through profits or dividends; labor takes its share through wages; governments take their share through taxes. Each must have a just share, and the question of what is a just share and the question of who-shall make the allocation are vital current political and economic issues. Certain political philosophies con- tend that government should take all the fruits and that some dictator or politician should have irrevocable discretion and power to distribute the fruits. Believers in democracy and free enterprise adhere to the view that capital is entitled to redemption and fair remuneration; that labor shall have a fair wage; and that governments shall make equitable levies. But since all human beings, both individually and in groups, are inherently selfish, there are many variable views in regard to the question of what ratios shall be applied in a division of the fruits and who shall have the power to fix these ratios. Perhaps all parties should devote more energy to means of increasing the crop, and not so much time to quarreling over a division of the existing crop. Moreover, all three parties must guard against killing the goose that lays the mineral eggs. For example, if the tax burden becomes oppressive and the taxpayer refuses to carry it, then business stagnates, with consequent unemployment, reduction of profits, and curtailment of government revenue. If a producer assumes the increased tax burden, marginal operations must be closed down, and again the consequences are increased unemployment and diminished government revenue. Quite often an increase in a rate of tax results in a reduction of government revenue. Perhaps the worst detrimental feature of selfishness and greed on the part of any one of the three distributees of the fruits of industry is the retardation of expansion of the enterprise. Growth means new jobs, more government revenue, better trade, and better standards of living. Explorations and discoveries have resulted in tremendous contributions

Jan 1, 1959

By Rustum Roy, E. F. Osborn

THE investigation discussed in this paper concerns phase equilibria in the alumina-silica-water system. Studies in this system are part of a re¬search project sponsored by the Geophysics Branch of the Office of Naval Research2 directed toward increasing knowledge of the stability relations of metamorphic mineral assemblages. Inasmuch as metamorphic rocks consist largely of aluminosilicate struc¬tures, many of them hydrated, the system alumina-silica-water is a logical starting point for the systematic laboratory study of metamorphic mineral systems. Although logical, this is not, however, an easy approach to the problem because of the very great difficulty of obtaining transformations among structures in response to changes in temperature, pressure, and con¬centration. Commonly equilibrium among the phases in this system is exceed¬ingly difficult to attain. New laboratory techniques are required to produce certain structural rearrangements, such as inversion of diaspore to boehmite or the transformation of alumina-silica mixtures to andalusite or kyanite. Phase equilibrium data for the alumina-silica-water system are consequently only partly complete, and only a summary progress report is presented here. Crystalline phases in the alumina-silica-water system are numerous and many are common minerals. The oxide components exist as: quartz, tridymite, cristobalite, corundum, water, and ice. Binary compounds occurring naturally are: sillimanite, andalusite, kyanite, and mullite in the alumina-silica system and gibbsite, boehmite, and diaspore in the alumina-water system. The most abundant among the ternary compounds is the composition A12O3.2SiO2.¬2H2O, existing in the four structural modifications, kaolinite, nacrite, dickite, and halloysite. The hydrated form of halloysite, A12O3.2SiO2.4H2O, is here called endellite. Pyrophyllite, A12O3.4SiO2.H2O, is the third ternary compound of natural occurrence. In addition to these minerals, two other ternary crystalline phases have been encountered in this study. The first is a structure resembling montmorillonite and corresponding perhaps to the pure aluminian end member of the beidellite series. The other has a composition approximating A12O3.2SiO2.H2O but its composition and structure are still in doubt. LABORATORY PROCEDURE In these laboratory studies, a starting material such as an alumina-silica gel or one or more of the crystalline phases is wrapped in platinum foil,

Jan 1, 1952

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

By H. R. Froning, R. O. Leach

In wertability alteration flooding, a chemical agent is rnoved through a reservoir by the flood water to increase oil recovery by decreasing the degree of wetting of the rock by the oil. Substantial amounts of the chemical may be lost during movement through the reservoir. The extent of the loss, and therefore the economics of the process, depends in some cases on factors which are difficult to reproduce in the laboratory. Therefore, a short-duration, low-cost field test method is needed to permit evaluation of chemical requirements under actual field conditions. This paper describes a small scale rest conducted at a single well for measuring chemical requirements, thereby giving a more reliable evaluation of this important factor in the applicability and economics of the process. In the rest a small wafer slug containing the chemical agent and a nonadsorbed tracer is displaced into the reservoir by a known volume of wafer. The well is then placed on producrion. Chemical loss per barrel of pore volume contacted is calculated from [he fractional recoveries of the agent tested and the nonadsorbed tracer. The method has been used to determine within the actual reservoirs the chemical requirements for both a sandstone and a dolomite reservoir. Several chemical agents are potentially available for wettability alteration flooding, although none is universally applicable. For some applications of the method, chemical costs per barrel of additional oil recovered can be substantially less than one dollar. INTRODUCTION Wettability alteration flooding provides a means of increasing oil recovery from reservoirs by decreasing the degree of wetting of the rock by the oil and increasing the displacement efficiency of the flood water. Earlier studies demonstrated a relationship between oil recovery during waterflooding and the degree of wetting of a rock surface by an oil. The application of wettability alteration flooding to the Harrisburg field of Nebraska provided a field test' of this recovery process. Subsequently, additional laboratory and field tests have developed additional procedures for evaluating wettability alteration flooding, and have indicated where the process may be applicable. Applicability of this process to specific reservoirs is determined by a progression of tests to determine sus- ceptibility of the reservoir to alteration of its wettability, to indicate the degree of recovery improvement and to estimate the amount of chemical required to process the reservoir. The economics of applying improved oil recovery processes depends not only upon the degree of improvement in oil recovery achievable by the process but also upon the process costs and the timing of the income and the investment. Emphasis in this paper is on the expenditure aspects of the process. The work reported in this paper indicates that the chemical investments required for wettability alteration flooding are substantial. For evaluating the economics of a potential flooding application it is imperative that a sound estimate of the chemical requirements be made for the reservoir. Generally, true reservoir conditions are not adequately simulated in laboratory chemical propagation tests. Because of wide well spacings, many years might be required to obtain chemical propagation data from conventional pilots or inter-well tests. Consequently, n short-duration, low-cost method is needed to determine chemical requirements in the field. The potential applicability of wettability alteration flooding is discussed, as well as the economics of wettability alteration with respect to the inherent and imposed restrictions on the timing of income and investments. DETERMINATION OF CHEMICAL REQUIREMENTS In the process of moving a chemical bank through reservoir rock, some of the chemical agent lags too far behind the flood front to be effective or is otherwise lost to the reservoir system. The extent to which these losses occur overshadows the reductions in chemical concentration due to diffusion and to mixing with the reservoir fluids. Experience indicates that almost without exception, chemicals which induce a wetting change undergo either sorption reactions or chemical reactions with mineral constituents of the pore surfaces. Other reactions may occur between the added chemical and the reservoir oil and water. Even in limiting consideration to reactions of the relatively inexpensive inorganic salts, bases and acids, the reactions may be exceedingly complex. Reservoir pore surfaces consist of more than silica in sandstone reservoirs, and more than calcite or dolomite in carbonate reservoirs. Many mineral species are present, each exhibiting specific tendencies to react with an injected chemical. The reactions which occur can consume enough of the agent to have an important effect on economics. These reactions can cause a change in pH of the chemical bank, or may remove some of the active chemical by precipitation, adsorption or reaction to form a new chemical which may or may not be effective in

By Ellis Clark

Within the last ten years the progressive party of China, headed by Li Hung Chang, the Viceroy of Chi-Li, has been making great efforts to develop the mining resources of that country, and particularly those that are within reasonable distances of the sea-board. Attention was first paid to the coal-deposits, especially those at KaiPing, which have been placed on a profitable basis with an output of 1500 tons per day, under the energetic management of Mr. Claude Kinder, M.I.C.E., and the successful financiering of Tong King Sing. These coal-mines were described in a report by Kwong Yung Kwang, edited as a paper for the Institute (Duns., xvi., 95.) The attention of the leaders of the progressive party has for the past five years been turned also toward the development of the precious and related metals—gold, silver, copper and lead. De-

Jan 1, 1891

By Wilson T. Lundy

The history of the production of sulfur from salt domes in Louisiana and Texas originated with the operations of the Union Sulphur CO. at Sulphur, La., followed by the Freeport Sulphur Co. at Bryanmound, Tex. Including the present development of Grande Ecaille by the Free-port Sulphur Co., which constitutes the tenth dome to be equipped for the production of sulfur in the Gulf Coast, less than 10 per cent of all the known domes have proved to contain sulfur in commercial quantities. The Grande Ecaille salt dome is in the tidal marsh of the Mississippi River delta in Plaquemines Parish, Louisiana. It is 10 miles southwest of the right descending bank of the river approximately 45 miles below the city of New Orleans and is within 4 miles of the shore line of the Gulf of Mexico. Terrain Characteristic of the delta region, the terrain consists of a low, flat, uninhabited area of marsh land intersected by many shallow lakes and bayous. With the exception of salt grasses, the region is devoid of vegetation and presents, in all directions, an unobstructed path to the vagaries of the winds. The general character of this area can be best visualized from the mechanism of its formation.' Offshore bars formed enclosing long narrow lagoons between the bar and the inner shore line, the lagoons communicating with the sea by means of tidal inlets. The water in the lagoons, being comparatively quiet, favored the deposition of fine debris composed of sand carried into the lagoons by action of the waves, tides and river-borne sediments from the land surface. As this material accumulated the bottoms of the lagoons were built up and eventually salt grasses obtained a foothold. The successions of growth and decay of these grasses formed a brown fibrous mat varying in depth from a few inches to several feet. This mat of vegetation virtually floats on an ooze that is several feet thick and is composed of alluvial sediments and very fine sand, and the whole is underlain by an unctuous clay. The decayed vegetation from the upper layer when dried has a peaty appearance.

Jan 1, 1934

By Warren B. Weeks

Continued development in the deeper fields discovered during 1937 and 1938 was largely responsible for the 16 per cent (2,913,400-bbl.) increase in production, bringing the 1939 output to 21,376,230 bbl., the largest for South Arkansas since 1929. The older settled fields continued to decline and Rodessa was added to this classification. Magnolia field accounted for much of the increase, with Schuler field showing only a slight increase. There were 248 wells drilled during 1939, five more than during the previous year. Of 198 wells drilled in proven fields, 21 were dry; 46 of the 50 wildcats drilled were dry. The old depth record of 8332 ft. was increased to the present record of 9028. Trend of Prospecting Production from the porous " Reynolds " oolite zone of the Smackover limestone continued to be the objective of prospecting. During the year 20 seismograph, 3 gravity meter, 1 magnetometer, 1 soil analysis and 1 electromagnetic crews were reported in the state. The geophysical work was largely concentrated in three southern counties: Columbia, Lafayette and Union. As indicated in Table 2, the same three counties received most of the important drilling tests. At the end of the year only three seismograph crews remained in the areas. The area for prospecting Smackover limestone was greatly narrowed during the year by the failure of two "fault zone" wells to find satisfactory permeability in the oolite zone, and by the relatively thin sections of porosity found in wells drilled in townships 18 and 19 South. Geology Most of the wells listed in Table 2 afforded additional control on structure of the lower beds, but gave little new information on stratigraphy. Of the 248 wells drilled, 3 ended in Tertiary, 34 in Gulf, 1 in Paluxy, 29 in Glen Rose, 15 in Travis Peak, 15 in Cotton Valley, 139 in Smackover, 2 in Eagle Mills and 10 in Carboniferous. The age of the beds between the Travis Peak formation and Carboniferous is still questionable. Most

Jan 1, 1940

By Warren B. Weeks

Continued development in the deeper fields discovered during 1937 and 1938 was largely responsible for the 16 per cent (2,913,400-bbl.) increase in production, bringing the 1939 output to 21,376,230 bbl., the largest for South Arkansas since 1929. The older settled fields continued to decline and Rodessa was added to this classification. Magnolia field accounted for much of the increase, with Schuler field showing only a slight increase. There were 248 wells drilled during 1939, five more than during the previous year. Of 198 wells drilled in proven fields, 21 were dry; 46 of the 50 wildcats drilled were dry. The old depth record of 8332 ft. was increased to the present record of 9028. Trend of Prospecting Production from the porous " Reynolds " oolite zone of the Smackover limestone continued to be the objective of prospecting. During the year 20 seismograph, 3 gravity meter, 1 magnetometer, 1 soil analysis and 1 electromagnetic crews were reported in the state. The geophysical work was largely concentrated in three southern counties: Columbia, Lafayette and Union. As indicated in Table 2, the same three counties received most of the important drilling tests. At the end of the year only three seismograph crews remained in the areas. The area for prospecting Smackover limestone was greatly narrowed during the year by the failure of two "fault zone" wells to find satisfactory permeability in the oolite zone, and by the relatively thin sections of porosity found in wells drilled in townships 18 and 19 South. Geology Most of the wells listed in Table 2 afforded additional control on structure of the lower beds, but gave little new information on stratigraphy. Of the 248 wells drilled, 3 ended in Tertiary, 34 in Gulf, 1 in Paluxy, 29 in Glen Rose, 15 in Travis Peak, 15 in Cotton Valley, 139 in Smackover, 2 in Eagle Mills and 10 in Carboniferous. The age of the beds between the Travis Peak formation and Carboniferous is still questionable. Most

Jan 1, 1940

By R. McNeill, D. E. Weiss, E. A. Swinton

In a continuous countercurrent exchange process, an alteration in any one of the operating conditions has a complex effect on the others, which can only be predicted by employing the transfer unit or the theoretical stage theory on a basis of trial and error. A simple method is described for illustrating diagrammatically the behavior of a counter-current system, the equations being simplified by means of a concept the maximum hypothetical exchange performance. An example based on a typical metallurgical system is given, in which a divalent metal is recovered from a dilute solution, the resin being regenerated continuously by a monovalent ion. Useful conclusions are drawn from a study of the theory. Practical methods for performing continuous ion exchange are discussed, and the development of equipment based on modified ore dressing jigs is described. A swinging sieve jig contactor is evaluated experimentally. DURING the last decade, the new synthetic ion exchange resins have been applied extensively in industries outside the field of water treatment, but there is no record of a continuous counter-current process operating on an industrial scale. Attempts have been made to devise a satisfactory process but many problems remain to be solved. The basic principles of continuous processes will be outlined, as well as the major problems in their operation and the progress made in the CSIRO laboratories toward the development of satisfactory industrial techniques. In the metallurgical field ion exchange resins can be used for various applications such as the recovery and concentration of valuable metals from mine waters,' the regeneration of pickling and plating liquors," the prevention of pollution by waste effluents and the recovery of the constituents from them," and the purification of valuable metals such as the rare earths by chromatographic fractionation on columns of ion exchange resins.7,8 . Turther applications undoubtedly will be found in the field of hydrometallurgy where the use of ion exchange resins would enable direct extraction of the desired metal ion from the filtered leach liquor or the leach pulp. For example, an ion exchange process has been described recently for the extraction of gold from a cyanide leach pulp." A continuous process would have advantages in many applications over the usual process employing a fixed bed and intermittent cycle. In a recovery process, it would yield a product stream of steady purity and concentration, it would waste less water in rinsing, and if the contacting apparatus were efficient less resin would be used, since each portion of the resin would be cycled as soon as it was loaded instead of lying idle until the whole bed was ready for regeneration. A very major advantage is that it would be simpler to control automatically. It is probable that continuous operation will be the key for really large scale applications of ion exchange. The flow sheet of a continuous ion exchange recovery-concentration process is illustrated diagrammatically in Fig. 1. Dilute liquor containing the valuable ion flows through the stripping section countercurrently to a moving bed of resin and leaves after a final contact with freshly regenerated resin. The resin leaves the unit almost in equilibrium with the incoming liquor and then flows to the regenerating unit where it is treated by a slow countercurrent flow of concentrated regenerant solution. The adsorbed ion is displaced from the resin and appears in the concentrated product stream. The resin then must pass through a rinse unit or section where regenerant entrained by the resin is washed back into the regeneration section by water. The regenerated and washed resin is then recycled back to the stripping section. I. Theoretical Operating Behavior of Continuous Ion Exchange Stripping System The simple theory of continuous ion exchange is analogous to that of solvent extraction and other diffusional transfer operations and is governed by the equilibrium relationship, the mass balance, the rates of mass transfer, and the contacting efficiency of the unit. Equilibrium Relationship—The relative affinity of two ions A and B, for a particular resin immersed in their solution, can be expressed by plotting compositions of the solution against compositions which exist in resin in equilibrium with those solutions, i.e. C/Co vs q/a where C, is the total normality of the solution, C is the normality of ion A in the solution, a is the total exchange capacity of the resin in gram equivalents

Jan 1, 1956