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By A. J. Fenn

Long experience with underground mining inspired the unusual idea of using an underground crusher-conveyor system as a substitute for surface haulage at the Cananea open-pit copper mine in Sonora, Mexico. Now implemented and working well, the system has born out the expectations of its originator, Cia. Minera de Cananea S.A. de C.V., that it would effect substantial cuts in the high cost of haulage. The Cananea deposit is a large, low-grade orebody above the mined-out Colorada pipe which was once one of the largest high-grade underground deposits in the district. Present facilities at the pit consist of: (1) A large open-pit mine capable of moving 80,000 tons of material per day with a shovel-truck, underground crusher and belt conveyor haulage system. (2) A 25,000-tpd concentrator equipped for four stages of crushing, rod mill-ball mill grinding with cyclone classification, and flotation in conventional machines. (3) A dump leaching process producing 1 million lb of copper per month in a tin-can scrap precipitation plant. (4) A two-reverberatory, six-converter smelter which can make 140 tons of blister copper per day. (5) A well-integrated contingent of supporting departments running the gauntlet from housing to power generation.

Jan 3, 1969

By R. G. Vardiman, M. R. Achter

THERE has been interest recently in the conditions under which dislocation pipe diffusion may be observed. In order to extend the measurements to temperatures below Tm/2, where pipe diffusion becomes more important, a chemical sectioning technique was developed for silver single crystals which makes possible the removal of layers of 100A or less. The procedure is briefly as follows. Silver single crystals of 99.995 pct purity were cut in the form of discs t by 4 in., electropolished, and electropolated with Ag110 to about 1000A thickness. The thickness was determined by calibrating the activity with the weight of heavier platings, up to 10,000A. Both diffused and undiffused samples were sectioned in a two-step process. Reaction with a solution of iodine in alcohol formed a thin, even layer of AgI, which was then dissolved in a KCN solution. The reaction rate of the iodine solution varied slightly over a period of time, so that the thickness of the stripped layers, while quite even and regular for a given specimen, ranged from 65 to l00W for different specimens. This variation was checked by using a thickly plated standard with each run. Layer thickness was determined from undiffused samples by the point where the layer activity drops to half the maximum. This maximum intensity also provides, with appropriate corrections, the initial concentration per layer co. Because of the small penetration distances, it is necessary to use the exact solution of Fick's law for a source of finite thickness.' For a plating of thick- where c is the concentration at depth x. flj the error function, D the apparent diffusion coefficient, and t the time. The diffusion coefficient was obtained from the concentration-penetration data by extrapolating to x = 0 where c/co = erf h/2(Dt)1/2 Enough points were taken to impart confidence in the accuracy of the intersection, generally about ten layers. Four temperatures were used as in Table I. For best accuracy, t and h were chosen to give values of the parameter (Dt)1/2/h between 1.0 and 1.7. A typical concentration-penetration curve is shown in Fig. 1. The diffusion coefficients calculated from Eq. [I] are plotted in Fig. 2. For comparison, the high-temperature data of Tomizuka and sonder 2 which go down to 630°C are plotted as the solid line 2(a) from their expression D = 0.40 exp (-44,090/RT). Also for comparison, the lowest temperature point of Hoffman and Tumbull3 is plotted. It is apparent that our data diverge appreciably from the high-temperature extrapolation by 373°C. A better straight line fit for the low-temperature portion is obtained by subtracting the extrapolated volume diffusion coefficient from the apparent. or measured coefficient, and plotting the difference, curve (h) of

Jan 1, 1970

By Behram H. Wadia, Robert S. Lehto

A new metallurgical treatment for capturing as much of the values as possible from zinc refinery feed has been devised by personnel in Cerro de Pasco's Peruvian operation at La Oroya. Results from a pilot test illustrate that 50 metric tons of throughput can be maintained whereby substantially all the silver and copper report in the sponge iron while zinc and lead with indium and cadmium volatilize in the fume. This new technique named the Zileret (zinc leach residue treatment) process complements the electrolytic system used in the refinery so that all values in a zinc concentrate are economically recovered. In the system, dried zinc leach residue is pelletized with soluble zinc sulfate which already exists in the leach residue and a finely divided reducing agent such as anthracite coal. Upon heat addition, the loose anthracite in the charge maintains a reducing atmosphere while the free space above the charge has an oxidizing nature. When the charge reaches the reaction zone at about 1100°C, metals such as zinc, lead, cadmium and indium volatilize from the pellets, are oxidized in the upper kiln atmosphere, and are then carried in suspension as an oxide fume to the cooler and baghouse. The copper and silver content in their reduced forms, along with the iron, discharge from the kiln as a calcined product which is then cooled to below 130°C before being exposed to the air.

Jan 9, 1969

By C. E. Birchenall, M. H. Davis

The rate of oxidation of titanium in the temperature range 650° to 950°C has been measured. 'The linear rate law obtained is explained by interface reaction control of the process. Tracer experiments indicate the oxygen diffuses in at least one phase of the scale. THE kinetics of oxidation of titanium, principally at low temperatures and under varying oxygen pressures, have been reported.¹,² In the present investigation the rates of oxidation of titanium have been determined by the gravimetric method in the temperature range 650" to 1000°C under 1 atm of oxygen. Some of the scales formed were powdered and an endeavor was made to identify the phases present by X-ray diffraction. Alexander and Pigeon1 ascribe a logarithmic rate law to the growth of thick films formed at temperatures up to 565"C, and Gulbransen and Andrew' suggest that the parabolic rate law holds up to 600°C. In both cases large initial deviations from those relationships are apparent and no explanation is offered. In the present experiments it is shown that a linear rate law holds in the temperature range 650" to 1000°C, and it is suggested that the initial deviations observed are due in part to a spurious surface area effect and also to the high solubility of oxygen in titanium metal. Positive information on the mechanism of oxidation is not available, but oxygen ion diffusion through the lattice is inferred from the re- sults of trace experiments. X-ray analysis of the scale confirms Gulbransen and Hickman's' observation on the presence of rutile TiO². Experimental Details The titanium used in these experiments was prepared by the reduction of titanium tetrachloride with magnesium and was used in the form of rolled sheet. Samples approximately 0.040x0.5x0.75 in. were cut and polished to a fineness of 4/0 on metal-lographic polishing papers. Individual samples, suspended on a platinum wire, were lowered into the hot zone of a furnace, in vacuo, by means of a winch, and oxygen was admitted after the desired temperature was attained. The increase of weight of single samples oxidized for fixed times at certain temperatures was determined by weighing the sample before and after oxidation on a chemical balance. In these instances the sample was used for only one time at one temperature. In some cases the course of oxidation was followed by attaching the sample to a sensitive spring balance and noting periodically the extension of the spring with time of oxidation. Kirkendall type experiments were carried out by marking the surface of the titanium prior to oxidation with a solution containing radioactive silver. After drying, the specimen was heated in vacuo to decompose the silver nitrate used, leaving a small amount of metallic radioactive silver on the surface. Following oxidation, these samples were mounted, ground, and then placed in intimate contact with

Jan 1, 1952

By J. F. Myers, F. M. Lewis

IN the mining and handling of sulphide ores, some degree of oxidation takes place on the sulphide surfaces, which are exposed to the atmosphere. It is, moreover, well known that the oxidation compounds formed are a factor when such ores are subsequently treated by flotation. Modern operating technique and reagents are capable of contending with this condition to a major degree, provided that the proportions of the reagents used are varied correctly. This is particularly true at the Tennessee Copper Co., where the ore treated contains pyrrhotite in addition to other sulphides. A number of soluble salts, some of which are unstable, can be formed during the oxidation of the sulphide ores. Those unstable salts, whose oxygen requirements are unsatisfied, are constantly changing to more stable forms because oxygen reacts with them during the crushing, grinding and other operations of ore preparation. We add no chemicals of any kind to the ore pulp in its preparation for flotation feed. Consequently the only compounds that are present in the flotation feed are those produced by the natural oxidation of the ore in the presence of air and water. The Tennessee Copper Co. ore is composed of from 60 to 70 per cent sulphides and from 40 to 30 per cent nonsulphide material. It is not the purpose of this paper to discuss the nature of these various oxygensulphur compounds nor their effect on flotation, but to describe a method for [ ] correcting the effect. Suffice it to say that we have found that the reducing sulphur and oxygen salts have a depressing action on our bulk flotation process.1 From long experience we find that the best means of correcting for this variation in our work is to increase or decrease the quantity of xanthate, the promoting agent, as the quantity of the reducing salts increases or decreases. It was, therefore, very desirable to provide the operators with some quick and

Jan 1, 1946

By D. C. Jones

THE term "coal-mine mechanization" is definable in several ways. In its general application it can mean the use of machines instead of manpower in each of the various mining activities necessary for the produc¬tion of coal-so that under this definition a mine that uses mechanical coal cutters could be classed as a mechanized mine. Common usage, however, insists that a mechanized mine is one in which machines are used for the entire process of mining or extraction and transportation of the coal. For the purpose of this paper, then, the second definition of "coal-mine mecha¬nization" will be used, and practices employed in mines where the greater part of the production is obtained through the use of mechanical loaders and conveyors will form the basis for this discussion. PHYSICAL CONTACT HAZARDS If we were to group all the hazards that accompany the use of mechanized equipment in a mine, the predominant class would be that which, for want of a better name, might be termed the "physical contact" hazards; that is, all the sources of danger that lead to accidents through physical contact of workmen with ma¬chines. A survey of accidents in a group of mechanized mines indicated that this class of hazards may be responsible for more than half the number of accidents in mechanized mining. For example, of 21 mining-machine accidents at one mine, 14 were of the physical contact type of 14 loading-machine accidents at the same mine, 8 were the result of physical contact. Certain of these hazards an controllable through the use of guards-¬a guard over the cutting-machine bits whet the machine is being trammed from place to place, or a chain or some similar locking device around the cutter chain, so that an accidental blow or push against the bit control will not cause the cutter chair to move. Other hazards, such as the possibility of. a loader operator having hi foot crushed by a crawler tread, or , brakeman being hit by a loader boon swung too far to one side while a car is being loaded, might be subject to control through the use of guard devices, but the reduction or elimination of all physical contact hazards by such methods doe: not appear feasible. Every accident of the physical contact type should be investigated to discover whether guards could have prevented it, but often the only way in which to combat these hazards will be through the awakened consciousness of each workman as to the danger that exists if he indulges in unsafe practices. A good approach to control of accident: in any mine is the establishment of a set o: safety rules or operating rules. The intent of such rules is to provide guides for the workmen to follow in the operation of equipment. Frequently the workmen have no regard for such rules, and observance is obtained only through compulsion or the part of management. An interesting development in connection with the formation of safety rules occurred during a training program for employees in the

Jan 1, 1946

By David J. Dunn, Richard G. Martin

Of many physical parameters critical to design of grinding processes, impact of grinding media is among the most difficult to measure or predict. Yet impact of falling grinding balls, pebbles, or rods accomplishes the fine grinding essential to metallurgical recovery of most important minerals. Unfortunately, the same impacts that break ore fragments deform or crack grinding mill liners. A project was undertaken at Climax Molybdenum Company to measure impact forces in operating ball mills so that they might be used to calculate stresses in liner materials and be correlated with service performance of mill liner alloys. Towards that end, a method of measuring impact forces on ball mill liners using an instrumented grinding ball has been developed. Preliminary tests show that impact forces and contact stresses imposed on bare mill liners are quite high but are greatly reduced by cushioning when coarse ore particles are crushed between balls and liners. Compressive and shear stresses induced in liners not protected by coarse ore are so severe that deformation or minute subsurface cracking are inevitable in all but the small production grinding mills, regardless of mill liner alloy strength or hardness.

Jan 1, 1979

By Eldon Dunlap

RECENTLY the producing branch of the petroleum industry has shown a considerable and growing interest in the quantitative determination of the water, oil, and gas content of cores as it relates to estimation of reserves, to providing a, basis for judging as to the feasibility of applying various secondary-recovery and pressure-maintenance processes, and in general to the operation of producing properties. For years it has been known that oil, gas and water coexisted in the pores of many semidepleted oil measures. It is now generally believed that oil-bearing beds contain connate water in amounts varying up to 50 per cent of the pore space. However, only lately has reliable direct evidence of the extensive occur-rence of connate water in oil sands been published. Very little work has been done to ascertain the effect of this water on the productivity and ultimate recovery of oil from wells. It was the intent of this research to study the influence of this indigenous water on the flow of oil through unconsolidated sands, with the hope that the results could be applied qualitatively to the flow of oil in underground reservoirs partly saturated with water. Obviously, ultimate recovery estimates based on the void volume of the reservoir will be in error unless a, conscious or unconscious correction is made for the volume of water occupying a portion of the effective pore volume of the reservoir- Evaluation engineers in the past have applied arbitrary saturation and recovery factors that may inadverently have compensated for the failure to consider specifically the water content. Nevertheless, the evidence available indicates that these factors need revising if each is to be independently correct and truly representative. Although the mere fact that the connate water renders part of the pore volume of the reservoir incapable of containing oil is of primary importance to the operator, it is also of interest to know whether and to what extent the presence of varying amounts of interstitial water influ-ences the permeability of the sand to oil.

Jan 1, 1938

By J. R. Long, E. T. Hayes, D. C. Root, R. S. Dean

THE U. S. Bureau of Mines has recently reported on the development of a process for preparing pure ductile titanium in substantial quantities' and on the physical properties that may be attained in the metal, when consolidated by powder metallurgy methods and fabricated by certain hot and cold-working procedures.2 The high physical properties (Table I), combined with excellent corrosion resistance and light weight, will lead no doubt to appropriate applications of the metal. In any application, methods of joining will be of considerable importance. The relatively high activity of the metal toward the common gases raises the question of weldability and conditions that might be required for successful welding. Since hydrogen, oxygen, and nitrogen embrittle the metal, it should be protected from these gases at elevated temperatures. This was demonstrated by a few preliminary trials of arc, gas, and resistance welding. In these trials only the resistance method proved to be immediately applicable. While atomic hydrogen welding of titanium may be successful in some instances, a considerable amount of development work will be required to prove its worth. In addition, it would be suited only to small assemblies that could be heated subsequently in a high vacuum to remove embrittling hydrogen. Heli-arc methods likewise would be limited to small assemblies, which could be completely protected by a helium atmosphere. It was quite apparent that titanium required more protection than was afforded by the normal flow of helium gas used for welding magnesium and aluminum. The hot metal is attacked by oxygen and nitrogen, and the rapid diffusion of the oxide into the metal prevents the formation of a protective coating and requires the helium atmosphere to be maintained until the weld has cooled nearly to room temperature. The preliminary work indicated that titanium could be successfully spot-welded without deterioration of the metal and that the microstructures of the welds were normal, exhibiting good bonding, and no oxidation. The high resistivity of titanium (53 microhms per cu. cm.) suggested that it would behave somewhat like stainless steel and allow considerable latitude in the several welding variables. The results reported here are intended to give a general picture of the effects of varying welding current, welding time, and tip pressure on single spot welds made on several different thicknesses of titanium sheet. The work was undertaken to demonstrate the feasibility of welding titanium rather than as an exhaustive study of spot welding. There are, therefore, a number of factors that will require further investigation to establish the specific controls required for actual applications.

Jan 1, 1946

Aalde, Kaare. ('40) Gen'I Del., Bisbee. Ariz. Aase, Glenn D., ('40) Univ. of North Dakota. Grand Forks. N. D. Abadesco, Enrique A., ('39) College of Engrg.. Univ. of the Philippines. Manila, P. I. Abbott. John L., ('40) Colorado School of Mines. Golden. Colo. Abdul-Ahad, A., ('39) Montana School of Mines. Residence Hall, Butte. Mont. Abreu, Ovidio M., ('38) Univ. of Nevada, Reno, Nev. Acomb, Allen. ('41) Louisiana State Univ., Box 1970, University. La. Acosta S., Gonzalo. ('40) Univ. of Texas, Univ. Sta., Box 1785. Austin, Tex. Adam, Howard W., ('41) Montana School of Mines. Butte, Mont. Adams, Albert, ('381 New Mexico School of Mines, Socorro. N. M. Adams. Charles W., ('39) Michigan College of Min. & Tech., Houghton, Mich. Adams. H. Travis, ('39) 316 South -5th St., Bur¬bank. Calif. Adams, Jack L., ('39) Univ. of Oklahoma, Nor¬man. Okla. Adams. John W., ('41) Colorado School of Mines, Golden. Colo. Adams. Lewis J.. ('39) 402 Del Mar Blvd., Corpus Christi, Tex. Adams, Rob L., ('40) Agricultural & Mech. Col¬ lege of Texas. Box 600, College Sta., Tex. Adamson. Charles R., ('41) Univ. of Utah, Salt Lake City. Utah.

Jan 1, 1941

By J. H. Bechtold, R. Kossowsky

1 HE superplastic behavior of the eutectoid A1-78 wt pct Zn alloy has been thoroughly investigated by Back-ofen et al.' The superplastic elongations and the marked resistance to necking were rationalized in terms of high strain rate sensitivity. In a subsequent paper2 superplastic behavior was formulated to show Nabarro-Herring diffusional creep as rate controlling at low strain rates, giving way to a dislocation climb process at higher strain rates. In a paper devoted to an interpretation of superplasticity in two-phase alloys, Packer and sherby3 have included the results of an examination of the structure of an Al-Zn eutectoid specimen before and after superplastic deformation. No significant change in grain size or directionality was observed in the deformed microstructures. This one observation led the authors3 to suggest that con- tinuous recrystallization or grain boundary migration was responsible for superplasticity. The purpose of this note is to report our observations on the superplastic behavior of the Al-Zn eutectoid alloy which differ in certain respects from works previously reported. Following Hayden et al., the results are discussed in terms of dislocation climb as the important factor during superplastic deformation. A heat of nominally high (99.995) purity 22 wt pct A1-78 wt pct Zn was induction-melted and cast into a 0.5 by 6 by 12 in. slab. The slab was heated to 350°C and rolled to 1/8-in.-thick plate. The plate was then cold-rolled to 0.040-in. thickness, annealed at 35OaC, and rolled to the final thickness of 0.020 in.* Speci-

Jan 1, 1969

By Edwin T. Myskowski, Alfred H. Hesse

THE development by the U. S. Bureau of Mines of a process for purifying manganese by electrolysis has stimulated interest in manganese alloys. While recent publications have shown some alloys of manganese and copper to possess rather unusual properties, none have included information on tensile properties, which have an important bearing on the usefulness of the alloys. Manganese-copper alloys can be heat-treated, since the single phases that exist at high temperatures may be retained by quenching. According to Persson,1 when quenched alloys containing from about 30 to nearly 100 per cent manganese are reheated, alpha manganese is precipitated (Fig. I). Calculations based on Persson's X-ray data indicate that this decomposition is accompanied by a change in volume of as much as 5 per cent. The reason that this large volume change has not been noted probably is the slow rate at which equilibrium is reached by these alloys. In an investigation on hardening by cold-work and heat-treatment, Dean and Anderson' found that quenched alloys increase in hardness with manganese content and that high-manganese alloys harden very slowly with cold-work, while high-copper alloys harden much more rapidly. Heat-treatment has little effect on manganese-copper alloys containing less than 40 per cent Mn, and affects those alloys containing more than 40 per cent in accordance with their manganese content and heat-treating temperature. Dean and Anderson4 also found that cold-work increased electrical resistance of alloys containing from 50 to 80 per cent Mn, and that heat-treatment of alloys containing from 40 to 96 per cent Mn altered the electrical resistance considerably. Dean, Anderson and Potter5 have shown that the high vibration damping capacity that characterizes manganese-copper alloys containing 40 to 90 per cent Mn is affected by cold-work and heat-treatments. Highest vibration damping capacity at low stresses was obtained by heating for 2 hr. at 450°C. (842°F.) after quenching from solution temperatures. Preparation of Alloys.-The alloys for these tests were melted in a 17-lb. basic-lined high-frequency induction furnace using electrolytic manganese, grade A ingot copper and a 70-30 Cu-Mn master alloy (69.80 per cent Cu, 0.12 Fe, 0.04 Si, 0.006 S, 0.07 C, balance Mn). The manganese was heated slowly to expel excess dissolved gases and melted before the addition of copper. No fluxes or covers were used. The melts were cast in tapered iron molds and the ingots were reduced 90 per cent when forged to ½ -in. rounds. These rods were given 2-hr. solution treatments at temperatures that depended upon the composition (Table I). Tensile Tests.-The tensile specimens, 0.313-in. diameter (1.25-in. gauge length), which were enclosed in evacuated Pyrex

Jan 1, 1942

By D. A. Dahlstrom

THE necessity for slime elimination from valuable mineral and coal products has become increasingly significant within the past 5 years.1,2 Most of the mechanized mining and present beneficiation methods have resulted in larger tonnages of extreme fines, silts, and clay, commonly called slimes, processed by the preparation plant. Economics in turn have forced the plant operator constantly to lower the bottom size of effective beneficiation, necessitating an almost complete recovery of valuable solids regardless of particle size. When these important factors are considered, the need for an efficient slime removal method for fine solids becomes self-evident. Ordinarily slimes have been classified as ranging from 75 microns, or 200 mesh, down to colloidal dimensions of less than one micron. At the same time there is general agreement that the most troublesome slimes are primarily clays, silts and extreme fines, so that the actual top size may be placed at approximately 30 to 40 microns. This immediately restricts any desliming operation to classification, as screening or other methods would be both uneconomical and grossly inefficient at such dimensions. The liquid-solid cyclone has been recognized as one of the most efficient fine size classifiers available to the mining industry. Applications of cyclones to thickening, high efficiency solid recovery, degritting,

Jan 1, 1952

By H. R. Stahl

The ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classi- fiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9½-in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 52¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1¾ in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges

Jan 1, 1947

By H. R. Stahl

The ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classi- fiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9½-in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 52¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1¾ in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges

Jan 1, 1947

By W. H. Howard

As the result of a series of analyses and volume determinations of gases discharged from the converters at the Garfield Smelting Co.'s smelter at Garfield, Utah, it was found that a considerable amount of lead fume was escaping from the stacks of the converter plant, and that if this lead could be recovered at a reasonable expense an attempt should be made to collect it. Two methods suggested themselves: (1) filtration through bags; (2) electrical precipitation of the condensed fume by a Cottrell plant. Filtration through Bags Tests were made in filtering converter gases through the ordinary woolen bags of lead-smelter bag houses to determine how long such material could withstand the corrosive action of the gases. While blowing on white metal only, the fabric was usually destroyed in 10 to 14 hr. In tests conducted in filtering gases when blowing on matte or slagging stage exclusively and when relatively considerable lead was being volatilizcd, bags in one instance lasted 60.5 hr. It therefore was evident that the Garfield converter gases did not contain in themselves sufficient neutralizing elements for the purpose of direct bag filtration, even during the slagging stage. Such results were more or less anticipated, but it was thus proved that under all conditions of blowing Garfield mattes a very considerable amount of neutralizing agent, zinc oxide or lime, or both combined, as in the Sprague process, would be required to render bag filtration possible. Again, owing to the constantly changing quantities of SO3 from a given battery of converters, depending on the number blowing at a given time and on the stages of the blow, it would be practically impossible to feed automatically the required amount of neutralizer. In order to safeguard the bags sufficient neutralizer would have to be continuously added to meet the maximum SO3 output, thus wasting neutralizers and giving an undesirably diluted final product. Such a condition is entirely different

Jan 1, 1915

By Ben L. Seegmiller

Introduction The profitability of an open pit operation depends to a large extent on the use of the steepest pit slopes possible, provided they do not fail during the life of the mine Optimum pit slopes can be designed by using rock mechanics technology. Such technology includes gathering and analyzing pertinent data prior to pit design, rigorous application of rock mechanics principles during the design process, remedial actions to improve slope stability, and monitoring slopes during the life of the mine. Ignoring rock mechanics technology in pit design may place the entire mining enter- prise in jeopardy Disruption of mining operations and safety hazards caused by slope instability can be severe. A classic case in point is the major slope failure (Seegmiller, 1972) that occurred at a southwestern US copper mining operation in 1970-1971 Minor slope stability problems initially amounted to no more than a nuisance, but in time, the unstable zones became larger, causing greater disruption of production and threatening safety Mining at or near the slide toe had continued throughout the period of slope failure, aggravating the already unstable condition The original plan for this pit called for mining the overall pit slope to 37.5" Mine operators were forced to lessen the back- slope angle in the hope that the failing rock mass would stabilize. However, stabilization did not take place, and when mining in the area of failure was completed, the overall slope had been reduced to approximately 27" Mining had been successful in removing only a small portion of the ore body at a high stripping ratio.

Jan 1, 1979

By L. B. Schmitt, N. H. Polakowski

Square bars of cold-dvawn copper, mild steel and a 2024 aluminum alloy were forced through an adjulstable gap between two hardened steel rollers by pulling or pushing. The rollers were either free to rotate or were locked in position so that conditions comparable with drawing, rolling, ad containerless extrusion could be created at will. Push or pull forces as well as roll separatirg forces were measured. Forming efficiency and coefficients of friction were determined. For- slow-rate deformation the values ot µ wrere between 0.09 and 0.14. Possibilities for further investigation are indicated. THIS work was undertaken with the aim of a) determining the principal parameters of cold drawing under simplified conditions directly, without recourse to speculative assumptions and b) comparing drawing with rolling and extrusion from the viewpoint of forces involved and deformation efficiency. Data of type a) relating forces in axially-symmet-rical drawing to die geometry. drafting schedules, material properties and lubrication conditions are abundant. Their intrinsic value is. however. limited as long as the radial die forces are unknown. On the other hand. the difficulties associated with measurement of these forces are serious as witnessed by the extreme paucity of such information."" Comparative investigations of type b) were also remarkably few and they all employed ring-shaped dies through which round bars were either drawn or pushed.2-4 Only in one case were the applied and the die splitting forces measured simultaneously.2 In order to simplify the problem, deformation under approximately biaxial conditions was adopted in the present work;. This enables both the drawing (rolling, extruding) efforts and the separating forces to be measured with reasonable ease by conventional methods. Since drawing, extrusion, and rolling can be carried out under virtually identical conditions only in a roller die. an adjustable tool of this type was constructed. The circular die profile involved a difficulty with interpreting the drawing and extrusion data. This was obviated by arbitrarily defining the angle between a contact chord and the drawing plane as the -.equivalent" die semiangle. The simplification was considered permissible since the effect of die angle was not specifically investigated. An additional conlplication was caused by the square cross section of the processed bars. This shape was selected as a compromise between the desire to have as massive as possible specimens and the capacity of the roller die. The metal was only partially restrained from spreading laterally. the flow conditions being therefore intermediate between plane stress and plane strain. On the Huber-Mises yield criterion the constrained and uni-axial yield strengths differ by up to 16 pct but the actual error due to the ignorance of the constraint factor was presumably much smaller than the maximum. In spite of these reservations it was felt that the method employed and results obtained have some merit. With certain further refinements to be indicated later, this technique can yield data perhaps more useful for checking metal forming theories than other, mostly indirect, methods. EXPERIMENTAL Materials—The metals used together with their principal mechanical properties are listed in Table I.

Jan 1, 1961

By John Blandy

EVERY new mining district has had its own peculiar experiences in inventing and experimenting upon new methods for the various operations of mining, and more particularly in the processes of crushing and dressing ores. As a matter of course, during this period many old things have been reinvented, patented, and cast aside, there to remain until at a future day other geniuses shall bring them forth again as new. In this respect the Copper Region of Michigan has not been behind other mining centres, and probably in no other part of the country has more money been expended in devising new machines and improvements upon old ones for the crushing of the rock. The appliances for washing the sand have not been so varied, simply for the reason that, having but one mineral, or rather metal, of high specific gravity to separate from rock material which varies but little in its character in any one mine, it requires much less care than is necessary in most mining centres of the world. At the beginning of operations in that district most of the work was in the control of Cornish miners, who introduced the simplest of Cornish mills, namely, wooden stem pestles, with wooden shafts and cams. These were well suited to the small mines, and particularly to the limited means of transporting more expensive machinery. As these facilities improved, they were enabled to change to iron, and to vary their patterns of rods, shafts, cams, and mortars. The most approved pattern, finally obtained, is the square or round stem, with collar adjustable by means of keys. In its present form it is a bar of cold rolled shafting, an eye in the top, an adjustable collar with key-plate and keys, the head and a shoe of chilled iron. The stem is fitted into the head by a slight taper. The battery has heretofore been of wood, lined with chilled cast plates, and bed-plate of the same, but within the past year the California pattern has been introduced. The screens are of sheet-steel, drilled with 16 holes to the inch. During this period of trial and gradual improvement in the pestle stamps, in the years 1855-6, the Ball steam-stamps were intro¬duced, and after years of labor and expense, have been made the most efficient and powerful machines ever yet used for the purpose. It is ostensibly the Nasmyth steam-hammer, and yet the

Jan 1, 1874

By Charles Hawes

WHEN sulfur is encountered in objectionable amounts, it is regarded as the most trouble-some element for the mine operator to control. It exists in two conditions in iron ores, as sulfide in iron pyrites and as sulfate in gypsum and barite. The sulfate radical (SO,) may also be linked to one or more of the several other elements found in the iron ore compound. The only dependable method for the de-termination of sulfur, particularly when found in the insoluble residue, has been the so-called fusion method. This is a long and tedious process involving several hours for its completion. In looking for a more rapid method which would be .applicable to all iron ores, the writer investigated the work of M. Groger1 who in 1881 showed that by heating pyrite with iron out of contact with air, the pyrite is reduced to iron sulfide and from the latter all of the sulfur will be given off as hydrogen sulfide gas on treat-ment with hydrochloric acid. FeS2 + Fe = 2 FeS FeS + 2 HCl = FeC12 + H2S In 1891 F. P. Treadwell2 not only confirmed the results of M. Groger but also found that iron powder would reduce to sulfide the sulfate radical of any substance, including barite, if not found in excessive quantities. The writer found this to be true and developed a method, in cooperation with F. J. Baker, chief chemist, mining department, of the Cleveland-Cliffs Iron Co., Ishpeming, Mich., for the determination of sulfur in iron ores which is not only very rapid but also accurate. Pure iron ore powder, reduced by hydrogen, gave good results but was found to be very slow in reacting. A high silicon iron with a low sulfur content of 0.012 per cent was found to be. much more active than the pure iron and gave excellent results.

Jan 11, 1927