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By John Wulff

IN a previous paper Burwell and Wulff1 have shown by electron diffraction studies that allotropic transformations can be induced in 18-8 stainless steel by polishing to a depth of about 5 X 10-5 cm. The development of a successful etching technique, which removes accountable amounts of metal from the surface without pitting and without changing the root-mean-square profile as measured with an Abbot profilometer, led to the application of the whole technique and material to an investigation of various mechanical methods of surface preparation, such as grinding, lapping, superfinishing and cold-rolling. Ideally suited to such a study is 18-8 of low carbon and nitrogen content. The material can be heat-treated to produce a coarse-grained single-phase structure. On cold deformation there occurs not only a reduction in grain size but, at regions of maximum stress, a change of face-centered material to body-centered material, provided the metal is worked below about 200°C. When worked between 200° and 415°C., diffraction studies evidence a reduction in grain size without the appearance of any new body-centered ferrite. Ferrite, which can be referred also to strains induced by carbide precipitation, is evident when the material is worked at temperatures between 415° and 800°C. Stainless steel is advantageous because the usual surface-finishing operations do not readily oxidize it to any appreciable depth. In order to produce detectable oxide films on the surface of austenitic 18-8 in a time equivalent to that required to produce, for example, a fine ground finish on ½ sq. in., the sample must be held in a furnace or oxygen stream above 600°C. By using longer heating times, it is possible to achieve measurable oxide films on samples whose surface temperatures exceed only about 250°C. EXPERIMENTAL PROCEDURE The apparatus, material, and technique described in the previous paper1 were employed in this investigation. Samples of 18-8, 1 by 1 cm. by 4 mm., were carefully annealed in the austenitic range to produce a coarse-grained structure. They were then quenched and pickled to remove scale. On a lapping machine of the Graton2 type, both sides of each sample were polished to a flatness of 1/6 a wave length of sodium D light. Thereafter they were electrolytically etched to remove at least 1 X 10-3 cm. of the surface, a depth that is well below that of any deformations induced by surface preparation. Samples of this material, carefully mounted in steel holders, were then sent to various industrial organizations known for special finishing techniques. In the instrument shops at the Massachusetts Institute of Technology check samples were finished by the same methods, as far as they were available. Abbot profilometer3 readings were made on all the samples, which were then carefully degreased in double distilled benzine and alcohol, dried in a hot-filtered air blast, and kept in a desiccator. Suc-

Jan 1, 1941

By L. P. Davidson

THE new electrolytic zinc plant of the American Zinc Company of Illinois commenced operation in April 1941. The simple flowsheet using the standard current density and the economic reasons that dictated it are given herewith. The Evans-Wallower Zinc Co. built an electrolytic zinc plant at Monsanto, Ill., in 1929 and .operated it for about two years. The plant stood idle from 1932 and was purchased by the American Zinc Co. in 1940. Because of the condition of the equipment, excepting the motor generator set, a complete rebuilding was necessary. The Fairmont City plant of the American Zinc Co. could treat the leach residues, the purification cake and the skimmings from the melting furnace in its various plants. This possibility permitted a simple flowsheet with good economic results. The flowsheet is shown in Fig. I. The calcined concentrates are leached in mechanically agitated tanks with spent electrolyte and such ferric iron solution as necessary. When the acid has been neutralized and the iron coagulated, the fluid is discharged to settling tanks. The underflow from these thickeners is filtered on a Moore filter; the filter cake leaves the plant for retreatment and the filtrate joins the thickener overflow. Zinc dust is added to this solution until the copper and cadmium are precipitated. The pulp is filtered through Shriver presses, which deliver the purified solution to storage tanks, and the copper-cadmium-zinc cake is shipped for retreatment and recovery of these three metals. The purified solution is electrolyzed; the regenerated acid is returned for use in the leaching plant and the cathode zinc, after stripping, goes to the melting furnace, from which it is cast in slabs. The skimmings from the melting furnace are shipped to the Fairmont City plant for reduction to metal in. the retort furnaces. The choice of this simple flowsheet was made possible only because of the facilities of the Fairmont City plant. The long leaching period, with high acid content and high temperature which characterize the high-density process, gives a higher zinc extraction but also dissolves more impurities, which must be removed. The same is true to a lesser degree of the multi-step leaching of the western low-density plants. By the sacrifice of some initial extraction, solutions can be obtained that are relatively easy to purify and give good ampere efficiency. This loss in initial extraction is made up in the retreatment of the residues in the Waelz plant. The impurities that accompany the zinc thus extracted are inconsequential when the Waelz product is treated in a distillation plant. It was recognized that this flowsheet would require a somewhat greater amount of sulphuric acid-either as acid or as sulphate sulphur in the calcine-because of poorer washing. The sulphuric acid plant at Fairmont City could deliver 60° Be. acid to compensate for the sulphate loss and the cost would be less than the cost of more intensive washing.

Jan 1, 1942

By L. P. Davidson

THE new electrolytic zinc plant of the American Zinc Company of Illinois commenced operation in April 1941. The simple flowsheet using the standard current density and the economic reasons that dictated it are given herewith. The Evans-Wallower Zinc Co. built an electrolytic zinc plant at Monsanto, Ill., in 1929 and operated it for about two years. The plant stood idle from 1932 and was purchased by the American Zinc Co. in 1940. Because of the condition of the equipment, excepting the motor generator set, a complete rebuilding was necessary. The Fairmont City plant of the American Zinc Co. could treat the leach residues, the purification cake and the skimmings from the melting furnace in its various plants. This possibility permitted a simple flowsheet with good economic results. The flowsheet is shown in Fig. I. The calcined concentrates are leached in mechanically agitated tanks with spent electrolyte and such ferric iron solution as necessary. When the acid has been neutralized and the iron coagulated, the fluid is discharged to settling tanks. The underflow from these thickeners is filtered on a Moore filter; the filter cake leaves the plant for retreatment and the filtrate joins the thickener overflow. Zinc dust is added to this solution until the copper and cadmium are precipitated. The pulp is filtered through Shriver presses, which deliver the purified solution to storage tanks, and the copper-cadmium-zinc cake is shipped for retreatment and recovery of these three metals. The purified solution is electrolyzed; the regenerated acid is returned for use in the leaching plant and the cathode zinc, after stripping, goes to the melting furnace, from which it is cast in slabs. The skimmings from the melting furnace are shipped to the Fairmont City plant for reduction to metal in the retort furnaces. The choice of this simple flowsheet was made possible only because of the facilities of the Fairmont City plant. The long leaching period, with high acid content and high temperature which characterize the high-density process, gives a higher zinc extraction but also dissolves more impurities, which must be removed. The same is true to a lesser degree of the multi-step leaching of the western low-density plants. By the sacrifice of some initial extraction, solutions can be obtained that are relatively easy to purify and give good ampere efficiency. This loss in initial extraction is made up in the retreatment of the residues in the Waelz plant. The impurities that accompany the zinc thus extracted are inconsequential when the Waelz product is treated in a distillation plant. It was recognized that this flowsheet would require a somewhat greater amount of sulphuric acid-either as acid or as sulphate sulphur in the calcine-because of poorer washing. The sulphuric acid plant at Fairmont City could deliver 60° Bé. acid to compensate for the sulphate loss and the cost would be less than the cost of more intensive washing.

Jan 1, 1942

By Basil B. Zavoico

The World War I1 spread to Russia on June 22, 1941, when the German Armies began their invasion. The Soviet industries, which were already engaged in a tremendous armament program, further accelerated the flow of munitions, and equipment and supplies for all purposes not directly military were either entirely eliminated or reduced to an absolute minimum. With the occupation by German armies of the Dniepr and Kharkov industrial regions, this trend was necessarily further accentuated. The oil industry of Russia, in consequence, was left to rely more and more upon itself, its own very limited manufacturing facilities, extensive repair shops and probably much used equipment out of order stacked in the warehouses. All evidence available to date points to the fact that the new situation not only did not affect the production either quantitatively or qualitatively, but, on the contrary, caused a larger flow of much needed fuels to the country's war machine, primarily by renaissance of both management and labor. The oil industry of the Soviet Union, left to depend entirely upon itself for equip ment and supplies, and even for new workers and technicians to replace those called into the services, did and is still doing a good and effective job. Since it could not hope for new equipment and parts and could not blame any other industry for the delays in delivery, it dragged out old machines from storage and repaired them in its own machine shops. Often such stored and used equipment was but slightly defective, waiting for some small part from factories in Central Russia, plants that now were either in German hands or entirely converted to the manufacture of munitions. More particularly, instead of the pipe that could not be taken to the Caucasus from Central Russia, casing reclaimed from abandoned or submarginal wells was used, or pipe from the abandoned or but slightly used pipe lines; and this practice is being continued even now on a large scale, to create an ample reserve for the duration of the conflict. Indirectly such a development was even an improvement on pre-war days, when many wells about to be completed had to wait for weeks, even months, for the arrival of new pipe, all blame then being placed outside of the industry on the factories in Central Russia. For the purpose of reclaiming equipment and pipe still usable, special crews were organized. The war also caused much quicker adaptation of certain American practices and introduction of many emergency measures, such as: removal of derricks upon completion of wells and the use of drilling equipment mounted on skids for rapid transfer from location to location—both of these practices, while standard in the U.S.A., being new in Russia; installation of simplified surface pumping units; substitution of wooden bearings in the surface pump equipment; completion of wells in the Ural-Volga district, where formations are hard and tight throughout the sedimentary column, without either casing or

Jan 1, 1942

By L. P. Davidson

The new electrolytic zinc plant of the American Zinc Company of Illinois commenced operation in April 1941. The simple flowsheet using the standard current density and the economic reasons that dictated it are given herewith. The Evans-Wallower Zinc Co. built an electrolytic zinc plant at Monsanto, 111., in 1929 and operated it for about two years. The plant stood idle from 1932 and was purchased by the American Zinc Co. in 1940. Because of the condition of the , equipment, excepting the motor generator set, a complete rebuilding was necessary. The Fairmont City plant of the American Zinc Co. could treat the leach residues, the purification cake and the skimmings from the melting furnace in its various plants. This possibility permitted a simple flowsheet with good economic results. The flowsheet is shown in Fig. I. The calcined concentrates are leached in mechanically agitated tanks with spent electrolyte and such ferric iron solution as necessary. When the acid has been neutralized and the iron coagulated, the fluid is discharged to settling tanks. The underflow from these thickeners is filtered on a Moore filter; the filter cake leaves the plant for retreatment and the filtrate joins the thickener overflow. Zinc dust is added to this solution until the copper and cadmium are precipitated. The pulp is filtered through shriver presses, which deliver the purified solution to stor- age tanks, and the copper-cadmium-zinc cake is shipped for retreatment snd recovery of these three metals. The purified solution is electrolyzed; the regenerated acid is returned for use in the leaching plant and the cathode zinc, after stripping, goes to the melting furnace, from which it is cast in slabs. The skimmings from the melting furnace are shipped to the Fairmont City plant for reduction to metal in the retort furnaces. The choice of this simple flowsheet was made possible only because of the facilities of the Fairmont City plant. The long leaching period, with high acid content and high temperature which characterize the high-density process, gives a higher zinc extraction but also dissolves more impurities, which must be removed. The same is true to a lesser degree of the multi-step leaching of the western low-density plants. By the sacrifice of some initial extraction, solutid'ns can be obtained that are relatively easy to purify and give good ampere efficiency. This loss in initial extraction is made up in the retreatment of the residues in the Waelz plant. The impurities that accompany the zinc thus extracted are inconsequential when the Waelz product is treated in a distillation plant. It was recognized that this flowsheet would require a somewhat greater amount of sulphuric acid—either as acid or as sulphate sulphur in the calcine—because of poorer washing. The sulphuric acid plant at Fairmont City could deliver 60" BC. acid to compensate for the sulphate loss and the cost would be less than the cost of more intensive washing.

Jan 1, 1944

By John V. Beall

Never having done it before, it took us all day and until 11 at night to select and pack for a four-day back- pack trip up Lake Chelan last month. When we were through, we couldn't lift our own pack and Sylvia's, OUT spouse's, we barely managed to budge. Nevertheless, the next morning we were at the dock to board the Laldy of the Lake for the 3 ½ -hr boat trip to Lucerne. Our dress and accouterments creating the appearance of a veritable expedition into the wilderness made a sharp contrast with our fellow passengers, a senior citizens tour group from Spokane. We had come out to Seattle for the SME Fall Meeting and were taking advantage of the occasion for a vacation trip back to Lucerne and Holden where we had lived briefly 25-years ago, and a backpack trip up Railroad Creek from Holden to the Glacier Peak country, weather permitting. No matter what you say, weather is important, and this is one of the ingredients that went into making the SME Fall Meeting and Exhibit in Seattle, September 22-24, a thoroughly delightful experience for the 700 plus persons who attended. Seattle is a beautiful city, but with late September sunshine, which we experienced, it is the best. The second ingredient, is a swinging host committee. Under the leadership of Don Anderson and Tom Van Zandt, co-chairmen, and Mrs. Cecilia Reed, Ladies Program, superb hospitality and fine arrangements were blended. Earl Beistline, from Fairbanks, organized a two-day short course on arctic operations which preceded the meeting and drew an attendance of 44. Dick Robbins, in charge of arrangements, provided Masters-at- arms for technical sessions and a presentation of the colors by the U.S. Coast Guard at the Welcoming Luncheon. Eric Cheney's field trips ranged far and wide with events both before and after the meeting. David (Jeff) Kroft, capably managed audio-visual operations with a group of students from the University of Washington. Charles Hochmuth worked long, hard and effectively before and during the meet- ing on finances. Hank Reed, managed an authentic three-ring circus as Entertainment Chairman. His activities covered two cocktail receptions for all registrants and a dinner-dance. Dolph Campbell and Joe Hackbarth provided publicity by press and TV (see p. 37-40). SME's meetings have certain characteristics and possibly not all of these are known to everyone. The first of these is the deliberate involvement of the AIME Section in the area of the meeting. Most trade shows and conventions are run 100% by a professional staff and although SME staff could also do this, it has been our experience that a more friendly and better meeting results when the local members apply their talents. Actually SME only considers a site for a meeting after it has received an invitation from a Section. The knowledge of local conditions which Section committeemen bring to a meeting is of great value.

Jan 1, 1971

By G. E. Archie

The usefulness of the electrical resistivity log in determining reservoir characteristics is governed largely by: (I) the accuracy with which the true resistivity of the formation can be determined; (2) the scope of detailed data concerning the relation of resistivity measurements to formation characteristics; (3) the available information concerning the conductivity of connate or formation waters; (4) the extent of geologic knowledge regarding probable changes in facies within given horizons, both vertically and laterally, particularly in relation to the resultant effect on the electrical properties of the reservoir. Simple examples are given in the following pages to illustrate the use of resistivity logs in the solution of some problems dealing with oil and gas reservoirs. From the available information, it is apparent that much care must be exercised in applying to more complicated cases the methods suggested. It should be remembered that the equations given are not precise and represent only approximate relationships. It is believed, however, that under favorable conditions their application falls within useful limits of accuracy. Introduction The electrical log has been used extensively in a qualitative way to correlate formations penetrated by the drill in the exploitation of oil and gas reservoirs and to provide some indication of reservoir content. However, its use in a quantitative way has been limited because of various factors that tend to obscure the significance of the electrical readings obtained. Some of these factors are the borehole size, . the resistivity of the mud in the borehole, the effect of invasion of the mud filtrate into the formation, the relation of the recorded thickness of beds to electrode spacing, the heterogeneity of geologic formations, the salinity or conductivity of connate water, and, perhaps of greatest importance, the lack of data indicating the relationship of the resistivity of a formation in situ to its character and fluid content. On the Gulf Coast it is found that the effects of the size of the borehole and the mud resistivity are generally of little importance, except when dealing with high formational resistivities or extremely low mud resistivities. Fortunately, little practical significance need be attached to the exact values of the higher resistivities recorded. Low mud resistivities are not common, but when this condition is encountered it may be corrected by replacing the mud column. With the present advanced knowledge of mud control, invasion of mud filtrate into sands can be minimized, thereby increasing the dependability of the electrical log. The effect of electrode spacing on the recorded thickness of a bed is often subject to compensation or can be sufficiently accounted for to provide an acceptable approximation of the true resistivity of the formation. As development of a field or area progressively enhances the knowledge of the lithologic section, the resistivity values of the electrical log take on greater significance, ultimately affording acceptable interpretations. The salinity, and

Jan 1, 1942

By John W. Bell

In the introduction to an excellent pamphlet, John Gross' makes the following statements: Although marked progress has been made along mechanical lines, the theory and conception of underlying principles have not advanced so rapidly. This lack of advance in theory may be attributed to the unfortunate situation resulting from the controversy as to whether the Rittinger or Kick law is applicable to crushing. Many pages of the technical literature are burdened with theoretical discussions in favor of one or the other of these laws, which have tended to cloud rather than to clear the atmosphere. Realizing the situation, the Milling Committee of the American Institute of Mining and Metallurgical Engineers, through its secretary, E. A. Hersam, studied the milling industry and, because of the apparently hopeless lack of agreement in the Kick versus Rittinger dispute (5, 13, 14, 16, 18. 19. 21, 22, 23, 24, 30, 34). concluded that a foundation lor the theory of crushing was of primary importance. The numbers in the paragraph quoted refer to papers listed by Mr. Gross in his Bibliography on Crushing and Grinding.I One is obliged to infer that these papers show "hopeless lack of agreement in the Kick versus Rittinger dispute." No. 22 is a paper by Arthur 0. Gates2 and No. 23 is one by the writer.3 Employing the subject matter contained in papers Nos. 22 and 232,3 the writer will show that their authors formulated identical conclusions, based solely on experimental evidence. Exactly the same defect in the Stadler hypothesis was described in each of the two papers, although entirely different experimental methods were employed in uncovering it. Both of the authors, independently, formulated the conclusion that Rittinger's was a true hypothesis and Stadler's false. The evidence in a third paper4 will be employed to prove that Stadler's hypothesis, when examined experimentally, is worthless. Stadler's Hypothesis Worthless It is the writer's opinion that the evidence available in 1915 and the corroborating evidence in 1916 was amply sufficient to eliminate the need for any further consideration of Stadler's hypothesis. Certainly the evidence produced in 1918 dealt this hypothesis its coup de grace. It follows, therefore, that in the year 1918 there was only one credible hypothesis left—the Rittinger hypothesis. Gates' Opinion What information is available in Mr. Gates' paper?2 His conclusions are not consecutively summarized, none the less they are perfectly clear. On page 876, Gates says, and it is a very significant utterance: But my results show, among other things. that many more of Stadler's energy units are obtained per foot-pound-applied in coarse than in finc crushing. This suggests a doubt as to the correctness of his unit, since one would expect a foot-pound-applied to produce the same number of "energy units," whether the feed and resulting product were coarse or fine. When Rittinger's theory is applied, the number of mesh-tons (my unit of surface produced) is

Jan 1, 1943

By S. F. Ravitz

In his excellent book on the Principles of Flotation, Warkl makes the following significant statement concerning the theory of flotation: Two questions of first-rate importance must be considered . . . First, it must be decided whether the sulfide minerals require a collector in order that they may float, or whether as has been claimed, they possess an inherent float-ability and need no collector. Second, it must be decided whether the xanthate type of collector can react with or be adsorbed by the unchanged sulfide minerals themselves, or whether, as many claim, reaction with the collector is dependent upon initial oxidation of the mineral surface. In the theory of flotation there are no two questions of greater importance than these. Ravitz and Porter2 presented evidence that galena can be floated without a collector, and even without any reagent, provided oxidation products are thoroughly removed from its surface. They found also that thoroughly cleaned galena floated faster when xanthate was used than when a frother alone or no reagent was used, and that, without a collector, the recovery decreased gradually upon exposure of the mineral to air. Their work, however, has met with considerable criticism.l(127).3-6 This paper is presented to answer the objections that have been raised, and to present additional data that confirin their results. The apparatus used to provide oxygen-free conditions for flotation has been described in detail by Ravitz and Porter. An atmosphere of purified nitrogen was maintained within the apparatus; all water and solutions used were freed of dissolved air, and could be transferred where desired by the use of nitrogen Pressure; nitrogen was used as the gaseous medium for flotation. The flotation cell was essentially a glass cylinder 15 cm. long and 4.6 cm. in diameter. Near the bottom was sealed a fritted glass filter, which broke up the gas into small bubbles, and 6 cm. above this filter was a side arm through which the froth overflowed. The pulp was agitated by a glass impeller, which entered the cell through a mercury seal. For each test 17 grams of galena and 75 ml. of water were placed in the cell, after which the speed of the impeller and the rate of gas flow were adjusted. Water was added gradually during the test to keep the froth level at the proper height. The mineral was prepared and cleaned in the following manner. Exceptionally Pure galena was crushed by hand on a clean bucking board, screened to —I50+ 200 mesh, and deslimed on a 200-mesh sieve with a fine spray of water. It was then boiled on a hot plate for a number of hours with several changes of the desired cleaning solution, after which both the solution and the mineral were placed in the "mineral flask" of the apparatus. The cleaning solution was drained off, and the mineral was boiled with oxygen-free water

Jan 1, 1943

By Fred W. Stiefel

This paper covers the operation and maintenance of Diesel locomotives underground on a portion of the Delaware River Aqueduct. † This part of the tunnel is 15 miles long, with shafts 14 ft. in diameter spaced about 5 miles apart. These shafts (2, 2a and 3) are 825 ft., 1551 and 840 ft. deep, respectively. The grade of the tunnel is almost level, except for the 12,100 ft. north of shaft 2, which has a grade of 2.31 per cent. The tunnel is circular, 17 to 19 ft. in diameter. It lies at a depth of 400 ft. below sea level, with a maximum overburden of 2400 feet. The tunnel driving was carried on in six headings, using 15 storage-battery locomotives, each of 13 gross tons and having a maximum speed of 8 miles per hour. The longest haul was 3 miles, which proved to be the maximum distance for which storage-battery locomotives could be employed economically for the amount of material to be moved. One locomotive was used for switching cars at each heading, one for switching cars at the bottom of each shaft and one on the main-line haul from the heading to the shaft. The maximum amount of material hoisted in one shaft from two headings when 104 ft. of tunnel was driven in 24 hr. was 1000 cu. yd., or about 2000 tons of rock in place. The total quantity excavated was 750,000 cu. yd. (1,500,000 tons of rock). The tunnel-lining operations using concrete presented a problem of hading a maximum distance of 5 miles in the tunnel. The concrete-placing equipment had a capacity of about 2000 CU. yd., or 4000 tons in 24 hours. Experience had shown that storage-battery equipment could not handle this quantity of material within the time required at the distance of 5 miles. Trolley-locomotive haulage was considered dangerous and expensive, and would cause interference because transmission equipment would be required. Therefore the performance of Diesel locomotives of 160 hp., with a speed of 18 miles per hour, in comparison with that of other types of locomotives, was studied. Choice of Locomotives In deciding upon the type of locomotive to be used for the 5-mile haul, four alternatives were considered, as follows: I. To purchase additional storage-battery locomotives with the necessary batteries over the 6 locomotives and 12 batteries available. 2. To convert the existing battery locomotives to the trolley type. 3. To purchase new trolley locomotives. 4. To purchase Diesel locomotives. Proposition No. i The 90-hp. storage-battery locomotives the (each weighing about 12 tons), when hauling a load of 64 tons, with a drawbar pull of about 1900 lb. on a level track, developed a speed of 8 miles per hour and could make five round trips in 8 hr. with a

Jan 1, 1943

By C. F. Seaman, R. G. Clay

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

By C. M. Marquardt

Tramp iron and steel moving on a conveyor belt cause small currents to be generated in a coil situated in a strong magnetic field, which are converted to an alternating current and are amplified. The output voltage from the amplifier fires a thyratron with a relay in its anode circuit, which actuates a howler and simultaneously drops a spot of marker material on the belt. THE problem of tramp iron removal from moving ore belts is a long standing one. When heavy ore streams are carried on a belt, magnetic pulleys and strong surface magnets fail to remove tramp buried in the ore stream. Tramp iron buried in the ore stream cannot be pulled from the bed by surface magnets and is carried past the magnetic pulley by the falling ore stream. With the increased use of detachable bits the problem of the detection and removal of tramp has become more necessary and difficult. Several types of tramp detectors have been developed. It would serve no practical purpose here to review exhaustively the literature on the subject. Methods of tramp detection used are: (1) The magnetic method, wherein the small current generated by the magnetized tramp passing a Coil is used. (2) The unbalanced oscillator method wherein the tank circuit of a stable radio- or audio-frequency oscillator is unbalanced causing a change of frequency or a change in plate current due to changes in eddy currents, hysteresis, or dielectric because of the presence of tramp. (3) The bridge methods wherein the impedance of one leg of an alternating current inductive bridge is changed due to the presence of conducting tramp. Each of these methods of tramp detection has its field of usefulness. The second and third methods are excellent for oxide iron ores, coal, sand and gravel, grain, etc. However, these latter methods are not suitable for use on ores that contain rich conducting sulphides, since a large piece of pyrite or galena will cause the same detector response as a piece of tramp iron. At the Caselton, Nevada, plant of the Combined Metals Reduction Co., sulphide ores are treated; therefore, it was necessary that a tramp detector operating on only the magnetic properties of the tramp iron be used. Tramp iron detectors of this type operate on the principle that if a piece of tramp iron moves near a coil in a high intensity, steady magnetic field, the presence of the moving tramp iron causes the magnetic flux lines through the coil to change. This induces a mall current in the Coil. In series with the Coil is generally placed a relay sensitive to currents of 1 to 3 microamp. When the tramp iron passes near the coil the flux linkages change, causing this very sensitive relay to actuate. Since the contacts of such a relay are very small it must actuate a second relay capable of breaking a larger current

Jan 1, 1951

By Raymond Mancha

A test of any fan requires the determination of such data as fan pressure, air volume handled by the fan, and power input to the fen shaft. When testing operating mine fans of the centrifugal type, test methods have been employed, for the sake of convenience, that permit great latitude in detail of procedure resulting in wide variance in results by different investigators. Some controversial aspects of field fan testing are the result of past failure to define clearly the various forms of fan pressure and to a lack of understanding of the accurate procedure for the measurement thereof. Also, the anemometer for air-volume measurement, although convenient, if carelessly used and maintained is subject to a wide error range with variation in method of application and in instrument accuracy. The appearance of the axial-flow type of mine fan offers the possibility of refinements in field testing that were impractical with the centrifugal fan. First of all, the fan inlet and outlet are more sharply defined, giving closer agreement as to just where the fan begins and ends; thus there is less disagreement as to the proper location for measurements of fan pressure. The area of passage into the inlet of an axial-flow mine fan usually constricts, causing high air velocities with accompanying steadiness of flow. Thus the fan inlet offers an ideal location for accurate air-volume measurements with the pitot tube, the proper use of which is well established by code. Thereby the controversial aspects of such measurements are eliminated. Use of the Pitot tube also facilitates a more accurate measure of fan pressure, since the individual pressures measured can be properly weighted, as will be explained later. Determination of power input to the axial-flow mine fan operating in the field is subject to the same inaccuracies as with the centrifugal fan if the axial-flow fan is also belt driven. However, because of the higher rates of rotation usual with axial-flow fans, direct-driven units are more numerous and it is unnecessary to assume a drive efficiency, thus eliminating one possible source of error. It is the purpose of this paper to define and explain the various types of pressure measurements involved in a field test of an axial-flow mine fan, and to further explain the proper use of the Pitot tube when measuring pressures and volumes. DefinitioN of Fan Pressures The two fan pressures most commonly discussed are fan total and fan static pressure, each of which has the same meaning regardless of whether the fan is blowing, exhausting, or blowing and exhausting, as in the case of a fan with ducts connected to both outlet and inlet. Fan total presscre is the amount by which the total pressure of the air delivered by the fan exceeds the total pressure of the air received by the fan. This pressure is a measure of the maximum amount of the work done on the air by the fan that is available for external use.

Jan 1, 1942

By R. W. Gurry

WHEN a steel in the austenitic state, with all its carbon in solution, is maintained, at constant temperature, in contact with a gas that removes the carbon from the surface, yet without otherwise altering the composition of the steel, the change in weight of the specimen, after any period of exposure, is a direct measure of the amount of carbon it has then lost. The purpose of the present paper is to show how, by application of established mathematical reasoning based on the fundamental diffusion law and on certain postulates, this loss of weight yields a complete picture of the distribution of carbon throughout a specimen of regular shape after any period of exposure, with all the accuracy required in any practical occurrence of decarburization. Precisely the same is true for the gain in weight during the reverse process of carburization, or for the gradual gain or loss of any element, so long as the steel comprises only a single phase at temperature, provided always that the experimental conditions conform to the postulates. The change in carbon content from surface to center of the specimen is always continuous, when everything is in solution, even though under the microscope the decarburized zone (or carburized rim) may appear to be sharply defined. The experimental procedure is described and some typical results are presented, which are then interpreted by application of the diffusion law. EXPERIMENTAL PROCEDURE Initially an endeavor was made to follow the gain of weight during the process of carburization in a hydrogen-toluene atmosphere, but it was soon found that the specimen became coated with graphite, which could not be removed without some loss of metal; therefore it was decided to continue carburization until the steel was saturated throughout, to remove all graphite from the surface of the specimen, and then to follow the loss of weight during decarburization in a moist hydrogen atmosphere at the same temperature as that used for carburizing. The carburizing procedure has been described in an earlier paper;1 it suffices to state here that the carburization lasted about 140 hr. at 760°C. (1760°F.) and 44 hr. at 1110°C. (2030°F.), these periods being long enough-on the basis of the calculations presented later-to ensure that the specimen was about 99.9 per cent saturated at each temperature. The specimen was a cylinder about 3.8 cm. long and 0.475 cm. in diameter; the materials were carbonyl iron, a commercial low-metalloid iron, and a steel with 3.6 per cent nickel and the usual proportion of other elements. The fully carburized specimen, after cooling at the top of the furnace tube, was placed in a lathe and sandpapered to remove all graphite; then the diameter was measured with a micrometer caliper, and the initial weighing (corresponding to complete saturation) was made. By means of a

Jan 1, 1942

By William B. Senseman

FoR reasons well known to mining engineers, wet grinding is quite universal in plants having to do with the extraction of metallic values from crude ores. In the processing of the nonmetallic and industrial minerals, dry grinding is the more common method. The crude materials usually contain less moisture than is required for wet grinding and the products usually are marketed in dry form; most of them as powders. To grind wet, therefore, would entail the addition of water and an expense for drying the product, which is not justified. Furthermore, the material would be a dried cake that would still have to be pulverized or disintegrated before use. For producing fine products, air separators have largely replaced the earlier practice of sizing with screens and bolting reels. Mills in which the air-separating feature is an integral part of the design are now commonplace. Mill designs are many and air systems have developed along various lines, but for the purpose of this discussion, only the fundamental idea need be emphasized— the idea of using air for sizing and removing pulverized material from a grinding mill. About 20 years ago, when the use of mills equipped with air separation had become common practice, heated air or gas was introduced to such a system for the removal of moisture. The purpose of this discussion is to outline the developments and trends that have followed this experiment. It is an important development, for mills equipped with air separation are now universally used for pulverizing not only nonmetallic and industrial minerals, but a great variety of other materials. All such mills are essentially the same. Dried material is pulverized, removed by an air stream, trapped in a cyclone collector, and relatively clean air from the cyclone is returned through a fan to the mill. There is no theoretical need for a vent from the system. The practical need is to vent the air that leaks into the system; and, since the cyclone is not 100 per cent efficient as a dust collector, this vented air carries some dust with it. The initial attempt at drying was made on a system like that described, by the addition of a furnace. The minus pressure within the mill draws hot gas from the furnace. Infiltration air, along with the drying gases and evaporated moisture, is vented as before. The first application was made to a mill used for pulverizing coal. The coal contained very little moisture, so very little drying effect was needed for the success of this experiment. Actually, some calculated efficiencies of over 100 per cent were obtained before due allowance was made for the conversion of mill power into useful heat. When moisture is low, the heat obtained from the conversion of milt power is an appreciable part of that required. That drying can be effected in the short time interval available in a grinding mill of this type is better understood by visual-

Jan 1, 1948

By W. Julian Parton

Treatment of the breaker wash-water discharges to remove the suspended particles of finely divided material is usually most efficiently and economically accomplished in settling or thickening tanks. The purpose of these tanks is to separate the excess water from the solids by sedimentation whereby the solid particles suspended in the water are thickened by the force of gravity. The water overflowing these tanks is either reused for coal cleaning or discharged directly to the streams. The settled solids, which are in a relatively concentrated condition, forming a sludge or thick pulp, can be disposed of by pumping to a storage dam, by slushing into mine workings, by processing for recovery of coal, or filtering by vacuum or centrifugal means and discharging with the breaker refuse. The scope of this paper will be limited to the discussion of settling and thickening equipment, the factors that influence the settlement of solids, and a brief description of tests that can be made to determine the size of settling facilities required for any particular condition. The practical application of sedimentation to the treatment of coal falls into two groups, one of classification or hydro-separation, and the other of clarification and thickening. Classification may be employed to prepare a relatively low-grade fuel, which can be used as is or can be further processed separately. Clarification and thickening is employed where the primary objective is to recover a relatively dewatered solid and return a satisfactorily clear water to breaker or streams. Operation and Types oF Settling Equipment When a stream of water flows across a tank, the time required for a particle to travel from the feed inlet to the overflow must be a little greater than that required for the finest particles or flocs in the feed to settle out of the stream. Theoretically, the liquid in a settling tank should rise at a more or less constant rate throughout the area of the tank Under these conditions the rising velocity of the water must be less than the settling rate of the finest particles if clarification is to be obtained. When classification of particles is to be made in the tank, the overflow velocity must be greater than the settling velocity of the particles that are to be removed with the tank overflow. The performance of a settling tank, then, depends on the settling rate of the solids and the current velocities in the tank. The oldest types of sedimentation employ intermittent types of settling tanks of any convenient shape. Means are provided for filling the tank with pulp and allowing it to settle until quiet, for moving the clean liquor and for removing the thickened sludge. The operation is necessarily intermittent but a sufficient number of tanks is employed so that a continuous flow of clear water is maintained. Continuous methods for the sedimentation of slurries were developed in the metallurgical industry. Cones from which settled solids were discharged by gravity

Jan 1, 1947

By J. D. Edwards

When Charles Martin Hall invented the electrolytic process for the production of aluminum, one basic requirement was a supply of pure alumina. Now, more than 50 years later, the same requirement still holds, and there are no technical straws in the wind to indicate that this situation is likely to change. Long ago, Aluminum Company of America took a number of steps to see that this country should not face shortages of alumina and aluminum in time of emergency. As one step to this end, the Company developed and used a supply of bauxite in South America in order to conserve dwindling Arkansas reserves. Another, and more a program than a step, was to investigate every type of process that might bear the possibility of utilizing the low-grade ores of aluminum that are more plentiful than bauxite. It is a long story, but the investigation included electric-furnace processes, acid extraction processes and alkaline extraction processes, with or without preliminary furnace and sinter treatments. Some of the processes developed were operated in commercial production. This work has been continuous since the first World War. The turn of the wheel came sooner than was expected when, in 1942, bauxite shipmerits through the caribbean became uncertain and hazardous through the operation of submarines. The tremendous requirements of aluminum for the war effort were consuming high-grade bauxite at an alarming rate and a national search for substitutes began. High-grade bauxites, for technical and economic reasons, have been defined as those containing about 7 per cent or less of silica. While bauxite of this grade was limited in amount in the United States, there was in sight a substantially greater volume of low-grade, high-silica bauxite not amenable to economic treatment by the Bayer process, but richer in alumina than the clays or other available low-grade ores. It was at this point, and after several years of intensive research on this particular phase of the problem, that Aluminum Research Laboratories brought to completion the invention of one of its research staff, which made practical the utilization of large reserves of low-grade bauxite in America. Bayer Process The Bayer process has been almost universally used for the production of alumina from bauxite. This process, in outline, comprises the digestion of hely ground bauxite in a solution of caustic soda (NaOH) under pressure and at an elevated temperature. Alumina in the bauxite goes into solution as sodium alumi-"ate and the impurities—oxides of iron, titanium and silicon—remain undissolved. The undissolved residue, known as "red mud,'' from its color, is separated and run to waste. AluKlinum hydrate (A12O3.3H2O) is precipitated from the solution of sodium aluminate, and the liquor containing the caustic soda and some alumina is returned to the Process for digesting more bauxite. The aluminum hydrate, after washing, is calcined at a high temperature to produce

Jan 1, 1949

By Walter S. Weeks

In this paper the effect of natural ventilation pressures is considered a part of the mine resistance and it is shown that their effect on the mine resistance is not the same as the total natural ventilation pressure of the shafts if the mine has several levels. It is shown that increasing the fan pressure causes the division of the air to approach the condition that would obtain if there were no natural ventilation pressure present. It is shown that if an appreciable difference in pressure is maintained between the beginning and end of a group of splits, the value of the integrated effect of the natural ventilation pressures on the individual splits does not change substantially with moderate changes of this pressure. A method is outlined for determining the resistance factor of the mine and the value of the effect of natural ventilation with the fan operating; also a method for determining this factor when circulation is caused by natural ventilation alone. Natural Ventilation Pressure Natural ventilation pressure (N.V.P.) is caused by the difference in weight of the air in two similar columns between two horizontal planes, and the difference in weight is caused principally by the difference in temperature of the two columns. The fan pressure causes some variation in density, but, since it is usually small and affects both columns, its effect is relatively unimportant. It has been suggested that "N.V.P." be called "stack effect"; but, since the term N.V.P. was used in a previous Institute paper,' applied in exactly the same manner, the term will be used. Frictional Resistance The frictional resistance to flow in a mine without natural ventilation may be expressed as RQ2 where R is a constant for a given air density and Q is the rate of flow. This frictional resistance plus the effect of velocity changes will be called the F pressure (F.P.) and a curve of F.P. against Q, the F characteristic of the mine. If a pressure is applied to a mine by a fan, the air flow so adjusts itself that the drop in pressure through any path from entrance to exit of mine, due to frictional resistances and velocity changes, is the same as that in any other path. The mine resistance varies as the square of the quantity. The air distribution without N.V.P. will be called "normal." Natural Ventilation Present When N.V.P. is irregularly distributed in a mine, it is more convenient in analysis to consider the N.V.P. as a part of the duct resistance. If it is so considered the resistance no longer varies as the square of the quantity, and if N.V.P. exists between levels the distribution of the air is no longer normal. The total N.V.P. of the mine is the pressure caused by the full length of the shafts. If a mine has several levels, N.V.P. exists

Jan 1, 1946

By W. E. Hess

AT the Vesta mines of Jones & Laughlin Steel Corp. on the Monongahela River, 35 miles south of Pittsburgh, JCM Joy continuous miners and 6-SC shuttle cars are used for pillar extraction in the Pittsburgh seam. The entire output of the mine goes into coke production for blast furnace operation. Because this coal, which has high metallurgical qualities, is close to the place where it is ultimately used and because supplies with these properties are being rapidly depleted, complete recovery is essential wherever possible. With the introduction of the continuous miner into the Vesta mines, therefore, it was necessary to plan a project which would facilitate development and provide for future total recovery. Use of continuous miners under varying conditions has necessitated changes with regard to roof action, method of development, method of pillar attack, and innovations in mining. In the following pages these changes will be discussed. Prior to the introduction of continuous miners the Pittsburgh seam at the Vesta mines averaged 72 in. of mineable coal, overlain with 12 to 16 in. of draw-slate, which weathers on exposure to air and is not firm enough to be held in place. Immediately overlying the drawslate is a wild or rooster coal, very irregular in thickness, varying from 12 to 30 in. Above the rooster coal the strata consists predominately of shales or lime shales and sandstone. Just before the introduction of continuous miners in the Vesta mines, however, a definite change took place in the character of the roof structure overlying the coal seam, see Fig. 1. Thickness of laminated roof immediately overlying the rooster coal varies with localities and introduces a problem in roof control. In fact, a set of eight main entries being developed in Vesta NO. 4 mine entered an area with so difficult a roof structure that management considered abandoning the section altogether. In this locality the roof consisted of thin alternate layers of roof coals and shales, with an occasional irregular seam of slick slate, the whole being laced with slips, and the coal itself being cut by clay veins and spars. In an effort to overcome some of the difficulties, roof bolting was introduced in 1949. This made it possible to recover coal from an area which otherwise would have been economically unmineable. There have been very few roof bolt failures. When conventional machines were replaced by continuous miners there were changes in the development of butt entries. By these changes it was hoped: 1—to speed development, 2—to provide as much solid coal as possible for the machine on retreat, 3—to avoid the necessity of crossing entries already driven, 4—to facilitate ventilation, 5—to provide a roadway for shuttle car haulage which would eliminate the necessity of running through doors or curtains, 6—to provide better roof control by increasing the distance between the center lines of entries, and 7—to provide more coal when the machine was in position to load.

Jan 1, 1956

By James J. Wise

INTRODUCTION One basic design feature utilized in all caving and sublevel stoping techniques is the drawpoint. These draw- points may be located on sublevels when using sublevel caving methods, or may be located at the floor pillar when using block caving, sublevel stoping, or vertical crater retreat methods. How the drawpoint is designed and developed, how the ore is withdrawn from it, and how the ore is hauled to the shaft or portal encompasses the loading and hauling function. Historically, predominate loading and hauling equipment for caving and sublevel stoping has been scraper slushes, air powered overshot loaders, and rail mounted train haulage. All of this equipment has two common constraints. 1. The equipment locked into specific work sites and is not easily shifted to another work site. This constraint can often lead to low utilization and low productivity of the equipment. 2. The equipment is designed to work in the horizontal plane, which requires that the various horizontally-oriented working levels be connected by vertical raises. Vertical development by means of raises is not only increasingly expensive, but also is not well adapted to movement of large equipment into the s topes. In the last 20 years, two revolutionary labor saving concepts have become common in caving and sublevel stoping. 1. Long-hole drilling, which has mechanized the development of the drawpoint extraction structures as well as the actual stoping and undercutting. 2. Vertical access by means of trackless ramps between levels, and the consequent use of rubber-tired vehicles not only for loading and hauling ore, but also for the rapid movement of men, material, and equipment into the stopes. The introduction of trackless mining systems for loading and hauling has not been easy. Mine planners have had to re-orient their thinking to the larger operating space required by trackless equipment. Larger access drifts have necessitated reconsideration of mine layouts to insure proper ground support. Spiral ramp developments were often laid out on slopes steeper than the equipment could negotiate economically. Mine operators had to accept the added responsibility of roadway maintenance, water removal, increased ventilation, and more technical mechanical maintenance. As the mine operators enthusiastically embraced the superior productivity of trackless mining systems, they some- times disregarded opportunities to combine the older types of equipment with trackless vehicles to maximize productivity at lowest cost. In order to gain a proper perspective on this problem, it is necessary to review the various techniques of loading and hauling and to become more aware of the advantages or disadvantages that each system or combination contributes. LOADING Loading consists of two functions, (1) primary loading from a drawpoint and (2) secondary loading into long- haul system, a truck, train, or possibly a belt conveyor. Mine planners have four basic loading systems available to fit various operating conditions. These are, (1) gravity, (2) scrapers, (3) rail-mounted loaders, and (4) trackless loaders. Each system can produce optimum results under specific conditions. Gravity Loading Gravity loading is probably the oldest system in general use today.

Jan 1, 1981