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By Elton A. Youngberg

The Holden mine operated by the Chelan Division of the Howe Sound Co. is on the east slope of the Cascade Range in north central Washington on the south slope of Railroad Creek valley at an elevation of 3500 ft. The mine may be reached by a 40 mile boat trip from the town of Chelan which is at the southern tip of Lake Chelan, to Lucerne at the mouth of Railroad Creek and an 11. mile bus ride up Railroad Creek to Holden. A11 freight and concentrate is moved over this route to Chelan Falls on the Columbia River which is on the railroad four miles below the town of Chelan. The mine is now producing 2000 tons of gold, copper, and zinc ore per day which is treated in the Holden mill. Gold-copper and zinc concentrates are made, the first of which is shipped to Tacoma, Wash., and the latter to Kellogg, Idaho, for smelting. Ore is broken by long-hole blasting using the Noranda system which has been modified to meet local conditions. Until recently, blastholes have been drilled by diamond drills. Now a partial substitution of percussion drill holes, drilled with tungsten carbide insert bits, is being made. Geology The ore body occurs as a replacement deposit in a highly metamorphosed series of sedimentary rocks, mainly gneiss and schists, in a shear zone several hundred feet in width and of undetermined length. Commercial ore has been found in mineable widths of 25 to 100 ft for approximately 2500 ft along its strike. The commercial minerals are chalcopyrite, sphalerite, and gold. During the period of mineralization considerable silicifica-tion took place giving the ore an abrasive drilling characteristic. Following the period of mineralization, numerous dikes were introduced into the ore body. The earlier ones were of granite composition having a width of a few inches up to 80 ft. These were followed by much younger, fine grained basic dikes which usually do not exceed 2 ft in width. Development of Percussion Blasthole Drilling Equipment Test work with the 1½-in. tungsten carbide bit was carried on in development headings for several months early in 1947. The short life of the bits, because of gauge loss caused by the abrasive nature of the rock, prevented its adoption for this use. However fast drilling speed and ability to drill a long uniform hole suggested its use for drilling blastholes in competition with diamond drills as diamond costs were steadily increasing and exper-ienced drillers were difficult to obtain. The 1½-in. bit was the largest available at the time initial test work was started with sectional steel. The 1½-in. hole limited the diameter of the steel thread and coupling which could be used. Type F couplings were first used but because of the small thread section excessive breakage of the steel was experienced. Type H couplings were tried next. In order to use this coupling which is 15/8 in. in outside diameter, it was reduced to 1 3/8 in. giving 1/8 in. clearance between the coupling and the hole. Rod breakage at the thread was substantially reduced but some coupling breakage was experienced, however the overall performance was considered satisfactory (see Fig 1 for illustration of coupling and thread). Early test work with the 1½ in. bit indicated machines of piston diameters larger than 255 in. would cause inserts to loosen or break. It was found however that the additional weight of the sectional steel cushioned the blow enough to prevent bit failures when 3-in. Leyners were used. Rods used with the 1½-in. bits were 7/8 in. q. o. for sectional steel and 1 in. q. o. for all chuck pieces. In May 1948, 2-in. tungsten carbide bits became available and test work was immediately started. The 2-in. hole approximated the AX (1 15/16 in.) diamond drill hole which was being used exclusively for blastholes and permitted their substitution for diamond drill holes in a ring without alteration of pattern, burden, or explosives. The 2-in. bit also gave room in the hole for larger couplings and permitted the use of heavier rods and 3½-in. machines, increasing the

Jan 1, 1950

By G. C. Rauch, T. H. Courtney, J. Wulff

The superconducting behavior of Nb-Ti alloys containing 40 and higher wt pct of Nb and variable oxygen content was studied as a function of thermomechanical processing. Critical current density (Jc) us applied transverse magnetic field (H) data were obtained for 0.010 in. diam wires at 4.2° Kin steady fields up to 120 kOe. The precipitate responsible for increase in Jc in each of the alloys studied was related to niobium and oxygen content. In alloys containing 40 to 60 wt pct Nb with less than about 1600 ppm of oxygen, w precipitation in the range 350° to 450°C appears to be responsible for the maximum value of Jc observed. In alloys of higher oxygen content (-2500 ppm), the oxygen-enriched a phase which precipitates in the temperature range 450° to 550°C is a more effective fluxoid pinner. Maximum Jc in low oxygen (350 ppm) 80 wt pct Nb alloys was achieved by 600°C aging heat treatments. Measurements of resistive critical field also reflect the changes in composition accompanying precipitation in these alloys. It is well known that the superconducting critical current density, Jc, of solid solution superconducting alloys can be increased by precipitation heat treatments.1-4 The precipitate, to serve as an effective superconzlucting fluxoid pinner and increase Jc, must be of proper size and spacing.' Its electronic properties must also be different from those of the matrix.6 As in age-hardenable systems, optimum temperature and time of aging are dictated by the composition and history of the particular solid-solution alloy under consideration. The alloy diagram of Ti-Nb7-9 shown in Fig. 1 is typical of ß-stabilized Ti alloys. Precipitation to achieve high Jc in Ti-Nb and Ti-Ta alloys is usually carried out in what is believed to be a two-phase a + ß region.1,2,4,10 Previous work on Ti-Nb4,11 and on other ß-stabilized systems such as Ti-V, Ti-Mo,12-14 has shown, however, that the first precipitate to appear is not necessarily (Y. Under specific conditions the meta-stable w phase is precipitated instead. A further complication arises in such alloys from the strong influence of contained oxygen on the phase stability and on the kinetics of precipitation.15,16 The superconducting critical current density of Nb-Ti alloy containing less than 36 wt pct Nb has been studied by Vetrano and Boom1 and Kramer and Rhodes.4 The most effective aging temperature was found to be about 425°C; fluxoid pinning was attributed to the precipita- tion of a by Vetrano and Boom and of w by Kramer and Rhodes. Although the latter workers used recrystal-lized material and the former cold-worked, studies carried out in our laboratory lead us to believe that the type of precipitate responsible for increased Jc is unaffected by prior cold-work.17 That other types of precipitates may also be effective, especially in higher oxygen-containing alloys, has been demonstrated by Reuter, et al.2 in this laboratory with the aid of electron microscopy. Reuter et al. found that at an aging temperature of 600°C, fee-TiO is precipitated, resulting in high values of Jc. Even higher values of Jc were observed after heat treatment at 500°C, but the precipitate could not be indexed as fee-TiO. Whether the precipitate was oxygen-rich w or a could not be clearly distinguished. The work nevertheless indicates that appreciable amounts of oxygen alter the nature of the phases precipitated at different temperatures, at least in alloys containing more than 36 wt pct Nb. Since such compositions are more readily worked into suitable wire than lesser niobium content alloys, it was decided to study the response to aging of the alloy compositions reported in the present paper with the hope of clarifying the effect of composition on the precipitation processes responsible for high values of Jc. I. EXPERIMENTAL Nb-Ti alloys containing 40, 50, 60, and 80 wt pct Nb were prepared by are-melting 35 g ingots of iodide crystal bar titanium and electron-beam melted niobium.

Jan 1, 1969

By George A. Laird

THIs old and once famous district played, through its enormous production of silver and gold, an important part in the history of the State of San Luis Potosi. According to a pamphlet prepared under the direction of the Governor of the State, Sr. Don Blas Escontria, and presented to the visiting members of the Institute during the Mexican meeting of 1901, this district is situated 20 km. east of the city of San Luis Potosi, in a range of mountains of moderate height, extending 135 km. N. and S., and 16 km. E. and W.; and the mineralized area hitherto developed is about 600,000 sq. m. The history of the district, as given by the record dated 1778, and formerly kept in the Sail Francisco convent, says that it was discovered in 1575 by Sr. Don Pedro Irarte, and adds that the city of San Luis Rey de Francisco de1 Potosi received its name (in allusion to the famous Potosi mines in Bolivia) by reason of the remarkable productiveness of its gold- and silver-mines in the first period of their exploration, during which they paid to the Crown, as quintos (the fifth part of the product), $97,000,000, as appears from the accounts of the royal treasury. The amount of the quintos paid during the " bonanza period " of 80 years was $72,000,000. The larger sum first mentioned probably represents a longer series of years. But either will suffice to indicate the great wealth of the mines. In 1663, the decadence of the industry began with the caving-in of several mines, concerning which the record of 1778, already cited, says, in substance, that on Good Friday one of the owners in the enterprise, in view of the unfavorable prospect of a lawsuit in which he was engaged, set fire to the timbers supporting the underground galleries and workings of the Sail

Jan 1, 1905

By Michael Hoch, Walter C. Wyder

The Cr-Ni system was investigated between 900" and 1400°C and 45 to 100 wt pct Cr. The Cr-Fe system was studied between 1200" and 1385°C and 70 to 100 wt pct Cr. Powdered alloy compacts were produced from high-purity chromium, nickel, and iron poulders. The compacts were sintered in an argon atmosphere and then machined into small, 0.030-in. diam cylindrical specimens. These were then heated by induction in the high-temperature X-ray diJfractwn camem under a helium atmosphere and irradiated with copper Ka radiation. From the results of the high-temperature X-ray diffraction patterns it can be concluded that chromium exists only in the bcc form, and that no eutectoid reaction occurs in the Cr-Ni system in the temperature interval 900" to 1400°C. The X-ray diffraction patterns are in agreement with the Cr-Ni phase diagram of Bechtoldt and Vacher, and Williams. It is possible that chromium undergoes one or more allotropic transformations between room temperature and its melting point. Most of the evidence was gathered from the study of the chromium-rich ends of the Cr-Ni and Cr-Fe binary systems. All the investigations rely heavily on indirect methods: thermal analysis, electrical resistivity, and micro-graphic inspection of quenched samples. In the present work the chromium-rich side of the Cr-Ni and Cr-Fe binary systems is studied, using high-temperature X-ray diffraction techniques. Thus the coexistent structures can be observed directly at temperature. PREVIOUS INVESTIGATIONS ~ansenl shows two possible Cr-Ni phase diagrams 1) a simple eutectic [Liquid =a-Cr(bcc) + y-Ni(fcc)], 2) a eutectic [Liquid+ P-Cr(fcc) + y-Ni(fcc)] and euctectoid [p-Cr(fcc) ==a-Cr(bcc) + yNi(fcc)]. The latter is based on the work of Grant and co-workers.2 Stein and rant'" placed the eutectoid composition at 1215°C and 68 wt pct Cr; their diagram is shown in Fig. l. Similarly, Sully3 also discusses the two proposed diagrams. Price and Grant,4 working on an isothermal section of the Cr-Ni- Fe ternary diagram at 1300°C, made no mention of a ß-chromium phase (fcc). The ß-chromium structure was also reported by Misencik 5 in the Cr-Nb(Cb) system. Williams 6 studied the Cr-Ni system using precipitation techniques. His work included the chromium and nickel solvus lines below 1250°C. He reported no ß structure nor any eutectoid reaction. The solvus lines which he produced were very similar to those of Jette, et al 7 below 1000°C. He concluded that Grant and coworkers2 had misinterpreted their data, and that effects which they had reported were actually due to the rapid and extensive precipitation of the nickel-rich phase. Bechtoldt and vacher8 used sintered powdered alloy compacts in an attempt to find the ß phase. They used metallographic and room-temperature X-ray diffraction techniques and also failed to detect the fcc chromium phase. They attributed the low slope of the chromium solvus at approximately 1150°C as the probable cause of the abundant precipitation obtained. They also stated that the precipitation was the probable cause for the abrupt discontinuities in the thermal analysis and electrical resistivity data which led to the conclusion that a eutectoid reaction existed. Fig. 1 also shows their diagram. Grigor'ev and coworkers9 investigated the polymorphic changes of chromium in the Cr-Ni phase diagram. Using standard microstructural and thermal analysis techniques, they published a phase diagram which contained five polymorphic forms of pure chromium and four eutectoid reactions. They stated that the room-temperature bcc chromium transformed to a fcc structure at 930°C, to bcc at 1300°C, to hcp at 1650oC, and then finally to bcc at 1830°C. The eutectoid reactions were placed at 850°, 960o, 1140°, and 1220°C. Also to be noted was the fact that the uppermost phase, the bcc above 1830°C, was different from the upper phase of Grant and coworkers,2 the fcc above 1840°C. Their eutectoid reactions, temperatures and compositions fall

Jan 1, 1963

By H. Rosenthal, A. L. Jamieson

THE action of mercury on stressed brass to produce cracks was known before Moore, Beckinsale and Mallinson1 showed that actual season cracking did not occur spontaneously but could be induced by ammonia. These investigators studied other substances, including diphenylamine, without finding that anything other than ammonia could cause season cracking. Grimston2 reported season cracking in cartridge cases stored in wooden boxes wetted with dilute sulphuric acid pickling solution. Season cracking associated with sulphur dioxide, water vapor and air has been reported by Johnston,3 although he reported that trimethylamine and pyridine did not cause cracking. Jevons4 has ascribed season cracking of brass in certain instances to trimethylamine, "aldehyde amine," "ketone amine" and pyridine, but no experimental work was done evidently to prove this definitely. Pyridine was investigated by Morris,' who found it to have considerable cracking power. In the present investigation, the object was a qualitative evaluation of a number of representative amines with respect to their ability to cause season cracking. Thus it was desired only that an appreciable vapor pressure of each amine be obtained in order to make the test as severe as possible, without attempting to test equivalent concentrations for comparative purposes. METHODS Specimens Two types of specimen were used: I. Unannealed 70-30 brass cups formed from 0.040-in. thick sheet. Height of cup was 1 3/8 in. and diameter was 1 7/8 inches. 2. First draw pieces of caliber .50 cartridge cases in the unannealed condition (70-30 brass). The cups were representative of a thin wall and thin base specimen, whereas the draw pieces have a thick base (1/4 in.) and a thick wall (approximately 3/32 in.). Both types of specimens contained high residual stress and cracked in less than one minute in a solution of I per cent mercurous nitrate and I per cent nitric acid provided that this was preceded by a 30-sec. pickle in 40 per cent nitric acid. Specimens were prepared as follows: (I) degreased in trichlorethylene, (2) rinsed in H20, (3) pickled 5 min. in 10 per cent by volume H2SO4 (1.84 sp. gr.), (4) rinsed in H2O, and (5) dried. Amines Tested In selecting the amines, representatives of the three series of amines and one series of substituted amines were chosen. The aliphatic series was represented by the methyl and ethyl primary, secondary and tertiary amines. These are the simplest aliphatic amines. For the aromatic series, the phenyl primary, secondary and tertiary amines were selected as being the simplest. The heterocyclic series was represented by pyridine, which is a simple heterocyclic base (tertiary amine).

Jan 1, 1944

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, 0, i8, 19, 21, 22, 23. 24, 30. 34), concluded that a foundation for 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.1 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 O. Gates2 and No. 23 is one by the writer.3 Employing the subject matter contained in papers Nos. 22 and 23 2,3 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 fine 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, 1942

By B. N. Bose, M. F. Hawk

SINCE Davenport and Bain' introduced the isothermal transformation technique for the study of austenite decomposition in steels, a new field of investigation has opened up. Extensive research has been carried out on many different steels, but relatively little work has been done on eutectoid transformations in analogous alloy systems. The researches of Smith and Lindlief,² Mack,³ and Klier and Grymko²0 on the beta-phase transformation in copper-aluminum alloys, of Hibbard, Eichelrnan, and Saunders4 and Smith5 on the kappa-phase transformation in copper-silicon alloys, of Chaudron and Forester6 on the decomposition of "FeO," and of Wasserman7 on similarities between the iron-carbon, copper-aluminum, copper-tin, and copper-beryllium systems essentially comprise the field. The iron-nitrogen system, in view of its remarkable similarity to the iron-carbon system, is particularly attractive for the study of eutectoid decomposition. The present investigation is concerned with the study of isothermal transformation, proceeding by the familiar nucleation and growth processes, and of martensitic transformation characteristics of pure iron-nitrogen alloys. The kinetics of transformation and the morphology of the transformation products have been investigated, for these must be the factors determining the properties and behavior of the final product. The Iron-Nitrogen System he binary iron-nitrogen phase diagram has been a matter of considerable disagreement among vari01.1s investigators. However, the phase diagram according to Eisenhut and Kaupp8 and Lehrer10 is now considered essentially correct as reproduced in fig. 1. The important values, 2.35 pct nitrogen and 591°C, for the co-ordinates of the eutectoid point (obtained from this diagram and used in the re- search to be described) were confirmed in a concurrent study of Paranjpe, Cohen, Bever and Floe." The eutectoid decomposition involves the transformation of the ? face-centered cubic interstitial solid solution of nitrogen in iron to the a body-centered cubic interstitial solid solution of nitrogen in iron and the ?' nitride of iron. This nitride is essentially the compound Fe,N; it also possesses a face-centered cubic lattice but a considerably larger one than the ?; the nitrogen atoms in ?' are arranged in an ordered way in the centers of the unit cubes.8,10,11 A martensitic structure has also been reported8,12 in iron-nitrogen alloys and was studied in the research to be described here. Materials and Methods The specimens used in this investigation were prepared by nitriding "pure" iron. This iron was supplied by the Westinghouse Research Laboratories and is the product bearing their trade name "Puron"; it contains about 0.04 wt pct total impurities, the major portion of which is oxygen. Specimens were ribbons several inches long, a tenth of an inch wide, and 0.125 mm thick; this was the

Jan 1, 1951

By John Wulff, J. H. Hollomon

The opinion has been widely held that the corrosion of alpha brass occurs by the selective solution of zinc. As late as 1939, Fink1 and Evans2 suggested that in the initial stage of the corrosion the selective solution of zinc produces copper-rich areas, which subsequently act as cathodes that facilitate the deposition of copper. Previously Abrams3 had shown that most of the copper found on corroded brass was re-deposited and not residual, although he did not consider the nature of the initiation of the corrosion process. Enough data exist to indicate that any metal surface is sufficiently heterogeneous to provide the anodes and cathodes necessary for the initiation of corrosion.4,5 Thus, for alpha brass we need postulate not an initial selective solution of zinc, but only the existence of local differences in electrochemical potential. The heterogeneity responsible for potential differences may arise from oxide and other films as well as from alloy concentration gradients. Even though selective corrosion may be possible in some alloy systems, previous data on alpha brass as well as the experimental work reported in this paper do not require this supposition. Bengough and May16 in a discussion of Abrams' paper, first suggested the addition of small amounts of arsenic (0.03 per cent) to a1pha brass to prevent the redeposition of copper (dezincification). This suggestion was later substantiated by Nixon.7 In their studies of the inhibiting influence of arsenic, Evans and Fink were led to the selective-solution hypothesis in order to explain the initial corrosion of such alloys. They believed that an arsenic film covered anode areas and prevented the formation of copper-rich spots capable of acting as cathodes for the deposition of copper. They did not consider, however, the mechanism by which the arsenical film was formed. Bengough and May,a in another publication, suggested that the commonly found gray or black arsenical film may act to: (I) carry oxygen to the brass surface, (2) increase the discharge potential of copper on brass, (3) precipitate cuprous chloride from colloidal suspension in chloride solutions. Masing9 assumed that the film found upon arsenical brass was metallic arsenic. His measurements of hydrogen overvoltage of bulk arsenic led him to believe that the inhibition of dezincification occurs because the overvoltage of a brass surface is lowered by the arsenic film. He reasoned that this film would permit the more rapid reduction of hydrogen at the cathode, which would prevent the redeposition of copper. In a recent paper,10 we have pointed out that the film formed during the corrosion of arsenical brass and copper in stagnant solutions is essentially arsenious oxide (As2O3),

Jan 1, 1942

By John Wulff, J. H. Hollomon

The opinion has been widely held that the corrosion of alpha brass occurs by the selective solution of zinc. As late as 1939, Fink1 and Evans2 suggested that in the initial stage of the corrosion the selective solution of zinc produces copper-rich areas, which subsequently act as cathodes that facilitate the deposition of copper. Previously Abrams3 had shown that most of the copper found on corroded brass was re-deposited and not residual, although he did not consider the nature of the initiation of the corrosion process. Enough data exist to indicate that any metal surface is sufficiently heterogeneous to provide the anodes and cathodes necessary for the initiation of corrosion.4,5 Thus, for alpha brass we need postulate not an initial selective solution of zinc, but only the existence of local differences in electrochemical potential. The heterogeneity responsible for potential differences may arise from oxide and other films as well as from alloy concentration gradients. Even though selective corrosion may be possible in some alloy systems, previous data on alpha brass as well as the experimental work reported in this paper do not require this supposition. Bengough and May16 in a discussion of Abrams' paper, first suggested the addition of small amounts of arsenic (0.03 per cent) to a1pha brass to prevent the redeposition of copper (dezincification). This suggestion was later substantiated by Nixon.7 In their studies of the inhibiting influence of arsenic, Evans and Fink were led to the selective-solution hypothesis in order to explain the initial corrosion of such alloys. They believed that an arsenic film covered anode areas and prevented the formation of copper-rich spots capable of acting as cathodes for the deposition of copper. They did not consider, however, the mechanism by which the arsenical film was formed. Bengough and May,a in another publication, suggested that the commonly found gray or black arsenical film may act to: (I) carry oxygen to the brass surface, (2) increase the discharge potential of copper on brass, (3) precipitate cuprous chloride from colloidal suspension in chloride solutions. Masing9 assumed that the film found upon arsenical brass was metallic arsenic. His measurements of hydrogen overvoltage of bulk arsenic led him to believe that the inhibition of dezincification occurs because the overvoltage of a brass surface is lowered by the arsenic film. He reasoned that this film would permit the more rapid reduction of hydrogen at the cathode, which would prevent the redeposition of copper. In a recent paper,10 we have pointed out that the film formed during the corrosion of arsenical brass and copper in stagnant solutions is essentially arsenious oxide (As2O3),

Jan 1, 1942

By W. L. Fink, L. A. Willey

DURING the war there was introduced a new higher strength aircraft alloy designated 75S.1,2.3 The properties of this alloy assure extensive applications in both military and commercial aircraft. It is therefore important to understand its limitations as well as advantages. The properties of 75S alloy in the heat-treated tempers, like those of many other aluminum alloys, are affected by the rate of quenching. In general, low quenching rates tend to cause low strengths and to adversely affect the resistance to corrosion. Commercial quenching procedures are usually sufficiently rapid to permit the development of maximum properties. However, in the center of heavy sections, such as large rolled and extruded shapes, the quenching rates may not be rapid enough to develop maximum strength or resistance to corrosion. The investigations of 75S alloy described herein reveal the most critical range of temperature and the effects of quenching sheet and extrusions at different rates through that temperature range, on the tensile properties and resistance to corrosion. PROCEDURE Preparation of Specimens Sheet specimens (3 in. by 9 in.) were prepared from a commercial lot of 75S "as rolled" sheet (0.064 in. thick). The composition° was as follows: Weight Percentage Cu Mn Mg Zn Cr Al with Normal Impurities 1.64 0.16 2.58 5.66 0.27 Remainder The specimens employed for the determination of the time-temperature relations during quenching of sheet were prepared by the following procedure: the hot junction of a 28-ga. iron-constantan thermoelement (asbestos insulated) was brazed to a thin strip of aluminum which was in turn brazed into the end of a 3/16 in. i.d. by ¼ -in. o.d. by 30 in. long aluminum protection tube. Before brazing, the tube was flattened to 0.064 in. thickness over the junction and adjacent lead wires for a distance of 1 ½ in. The flattened end of the tube was fitted into a slot cut from one edge to the center of the 3 in. by 9 in. by 0.064 in. sheet and brazed into place. The brazed joint was dressed to the thickness of the sheet and the entire specimen buffed to approximately the brightness of "as rolled" material. Because of the brazing operations required, these specimens were prepared using 3S alloy. For the investigation of quenching rates in heavy sections, there were used 75S extrusions which had compositions similar to that of the sheet. The specimens consisted of 14 in. lengths of ¾ in., 2 in., 3 in. and 4 in. diam cylindrical rods and a roughly rectangular extruded section which was approximately 4 in. by 7 ½ in. Thermocouples were peened into the bottoms of 1/8 in. diam wells drilled into one end and parallel to the axis to a depth of

Jan 1, 1947

By H. L. Gilbert, C. Travis Anderson, W. J. Kroll

IN a previous paper,1 the use of a split graphite tube resistor as a heater element for high-temperature furnaces has been described. The principal advantages of this type of construction are: I. The quadruple resistance or voltage compared with the top-bottom clamp resistor furnace. 2. Free expansion of the resistor. 3. The ease of protection from oxidation by air. 4. The firmness by which it may be held in the compact split clamp. This construction appeared to be well adapted for the fusion of high melting metals which require fusion in a vacuum. Zirconium metal had previously been melted at the Albany Laboratory of the Bureau of Mines in a vacuum arc furnace on a water-cooled plate, using a movable tungsten top electrode.2 When larger furnace capacities were desired, this type of construction proved to have its limitations as the largest tungsten rods available were about 7/8in. in diameter. The supply of direct current of high amperage for producing the . desired arc characteristics was also not available. Considerable difficulty was also found in compensating the atmospheric pressure on the flexible tubing of the movable electrode, at the same time keeping its movement sufficiently free to enable the striking of an arc on any part of the metal briquet. Among the various materials used for the electrodes in conducting experiments in the vacuum arc furnace was graphite. It was found that zirconium fused in a graphite cup with a carbon arc was contaminated to only 0.2 pct carbon. It was therefore decided to melt zirconium directly in graphite crucibles. A high-frequency furnace was first used in the attempt to melt the zirconium in graphite crucibles. The high-frequency heating coil was placed outside a silica tube which was evacuated by means of an oil diffusion pump backed up by a mechanical pump. The graphite crucible was thoroughly degassed at 1900°C before use. Considerable difficulty was experienced in operating at such temperatures in vacuo because, as shown by Zintl,3 carbon vapor formed and reacted with silica according to the reaction: Si02 + C = SiO + CO. The silicon monoxide which is volatile at these high temperatures reacted with the zirconium, thereby contaminating it with silicon and oxygen. The presence of carbon monoxide was detected by means of a gas discharge tube placed in the vacuum system as indicated by the pale blue glow of the tube as soon as carbon vapor started to react with the hot silica. The reaction between the carbon vapor and hot silica was avoided by the use of split cylindrical molybdenum shields which were inserted between the graphite crucible

Jan 1, 1948

By C. E. Lesher

THIS paper reports a study of the reactivity of 950°F and 1650°F cokes as measured by relative rates of reduction of iron oxides at temperatures up to 2200°F. Previous work cited shows general acceptance of the theory that reduction by carbon is a gaseous reaction, and that kind and character of carbon as well as particle size have measurable effect on the velocity of reaction. As will be shown, the data obtained in this study confirm those conclusions. The work was not designed to examine iron oxide reduction equilibrium, but if reaction velocity be defined as the speed with which "a reaction tends to approach conditions of equilibrium," the data here presented may be considered as a study of reaction rates, and the relative degree of reduction to metallic iron as the measure of reactivity. Three standardized combinations of Lake Superior brown iron ore with carbon were tested by similar procedures. One combination was a mechanical mixture of carefully sized high-temperature coke (1650°F) with the ore. The second was a mechanical mixture of the ore with Disco* obtained by carbonizing the identical coal at 950 °F. The third was an agglomerate prepared by carbonizing the coal and ore at 950°F, premixed in proportions to give as nearly as possible the same relative amounts of carbon and ore as the mechanical mixtures. This agglomerate, obtained by heating the finely divided ore (through 30 mesh) with coking coal through the plastic temperature range so as to form solid aggregates, gives a product in which the oxide particles are impregnated with, and intimately bound together with low-temperature coke. The agglomerate-—ore-Disco—was most active in oxide reduction; the mechanical mixtures of Disco and ore next in order, with coke the least reactive. General Discussion: Carbon exists in many forms and it is well known that the form or nature of the carbon used in reduction of oxides is related to the critical temperature of reduction. Sugar carbon, charcoal, and lampblack are forms of carbon that will reduce oxides at lower temperatures than high-temperature coke, and coke will, in turn, give a lower critical reduction temperature than graphite. There have been many investigations of this characteristic of carbons. Johnson' reported a difference of 130°F (70°C) in the critical reduction temperature of zinc oxide as between charcoal 1891 °F (1033°C) and Acheson graphite turnings 2021°F (1105°C) with zinc oxide. Bodenstein2 using charcoal and coke, found a difference of 138°F (77°C) comparing an experimental figure of 2066°F (1130°C) for coke and 1928°F (1053°C) for charcoal, in the reduction of zinc oxide. He concluded that this is very marked and observed that the "type of carbon merely raises or lowers the temperature at which rapid reaction takes place." Comparing the effectiveness of types of carbon in reduction of zinc oxide, it was found that a "brown coal coke" gave 97 pct zinc elimination at 1832°F (1000°C), as compared with 48 pct with "hard coal coke."' A wide range of metallic oxides was studied by Tammann and Sworykin,4 who found that the temperature at which decomposition of oxides begins depends on the nature of the carbon used. Carbon in the form of graphite, lampblack, and sugar carbon was investigated. Sugar charcoal will reduce Fe2O3 to Fe3O4 at 842°F (450°C) as compared with 1112°F (600°C) for coke, according to Meyer." Direct reduction of iron oxides by charcoal begins at 1382°F (750°C), but "first becomes intense" at 1652°F (900°C), whereas with coke, direct reduction begins at 1742°F (950°C), and "first becomes appreciable" at 2012°F (1100°C).6 he total reduction of the sample under certain conditions when heated in a current of CO with charcoal was about 100 pct for limonite and about 77 pct for magnetite. Using coke under the same conditions, the respective percentages were 75 and 47. In a study of processes for sponge iron7 by the Bureau of Mines, the conclusion was reached that a low-temperature char from noncoking subbituminous coal is the most satisfactory solid reducing agent. In a critical study of zinc smelting from a theoretical viewpoint Maier8 concluded that the reduction is by CO, that the reaction between ZnO and CO is intrinsically more rapid than the subsequent reduction of CO2 by C, which is limited by diffusion rates, which in part effectively limits the smelting process. Maier said that the operation is improved with the activity of the reducing carbon. An active carbon, he said, is one maintaining a low CO, content in the retort. Reactivity of Carbon: One form of carbon is more potent in reducing oxides than another. A carbon that reacts faster than another at a given temperature is said to be more reactive. Reactivity is measured by several methods, using carbon dioxide, air, or steam as reactants.9 ebastian and Mayers" have developed a method for the determination of absolute reaction rates between coke and oxygen by a study of ignition points under certain conditions. These and other investigators have established the relative reactivity of types of carbon. Lignite, charcoal, bituminous coal, cokes in the ascending order

Jan 1, 1951

By C. E. Lesher

THIS paper reports a study of the reactivity of 950°F and 1650°F cokes as measured by relative rates of reduction of iron oxides at temperatures up to 2200°F. Previous work cited shows general acceptance of the theory that reduction by carbon is a gaseous reaction, and that kind and character of carbon as well as particle size have measurable effect on the velocity of reaction. As will be shown, the data obtained in this study confirm those conclusions. The work was not designed to examine iron oxide reduction equilibrium, but if reaction velocity be defined as the speed with which "a reaction tends to approach conditions of equilibrium," the data here presented may be considered as a study of reaction rates, and the relative degree of reduction to metallic iron as the measure of reactivity. Three standardized combinations of Lake Superior brown iron ore with carbon were tested by similar procedures. One combination was a mechanical mixture of carefully sized high-temperature coke (1650°F) with the ore. The second was a mechanical mixture of the ore with Disco* obtained by carbonizing the identical coal at 950 °F. The third was an agglomerate prepared by carbonizing the coal and ore at 950°F, premixed in proportions to give as nearly as possible the same relative amounts of carbon and ore as the mechanical mixtures. This agglomerate, obtained by heating the finely divided ore (through 30 mesh) with coking coal through the plastic temperature range so as to form solid aggregates, gives a product in which the oxide particles are impregnated with, and intimately bound together with low-temperature coke. The agglomerate-—ore-Disco—was most active in oxide reduction; the mechanical mixtures of Disco and ore next in order, with coke the least reactive. General Discussion: Carbon exists in many forms and it is well known that the form or nature of the carbon used in reduction of oxides is related to the critical temperature of reduction. Sugar carbon, charcoal, and lampblack are forms of carbon that will reduce oxides at lower temperatures than high-temperature coke, and coke will, in turn, give a lower critical reduction temperature than graphite. There have been many investigations of this characteristic of carbons. Johnson' reported a difference of 130°F (70°C) in the critical reduction temperature of zinc oxide as between charcoal 1891 °F (1033°C) and Acheson graphite turnings 2021°F (1105°C) with zinc oxide. Bodenstein2 using charcoal and coke, found a difference of 138°F (77°C) comparing an experimental figure of 2066°F (1130°C) for coke and 1928°F (1053°C) for charcoal, in the reduction of zinc oxide. He concluded that this is very marked and observed that the "type of carbon merely raises or lowers the temperature at which rapid reaction takes place." Comparing the effectiveness of types of carbon in reduction of zinc oxide, it was found that a "brown coal coke" gave 97 pct zinc elimination at 1832°F (1000°C), as compared with 48 pct with "hard coal coke."' A wide range of metallic oxides was studied by Tammann and Sworykin,4 who found that the temperature at which decomposition of oxides begins depends on the nature of the carbon used. Carbon in the form of graphite, lampblack, and sugar carbon was investigated. Sugar charcoal will reduce Fe2O3 to Fe3O4 at 842°F (450°C) as compared with 1112°F (600°C) for coke, according to Meyer." Direct reduction of iron oxides by charcoal begins at 1382°F (750°C), but "first becomes intense" at 1652°F (900°C), whereas with coke, direct reduction begins at 1742°F (950°C), and "first becomes appreciable" at 2012°F (1100°C).6 he total reduction of the sample under certain conditions when heated in a current of CO with charcoal was about 100 pct for limonite and about 77 pct for magnetite. Using coke under the same conditions, the respective percentages were 75 and 47. In a study of processes for sponge iron7 by the Bureau of Mines, the conclusion was reached that a low-temperature char from noncoking subbituminous coal is the most satisfactory solid reducing agent. In a critical study of zinc smelting from a theoretical viewpoint Maier8 concluded that the reduction is by CO, that the reaction between ZnO and CO is intrinsically more rapid than the subsequent reduction of CO2 by C, which is limited by diffusion rates, which in part effectively limits the smelting process. Maier said that the operation is improved with the activity of the reducing carbon. An active carbon, he said, is one maintaining a low CO, content in the retort. Reactivity of Carbon: One form of carbon is more potent in reducing oxides than another. A carbon that reacts faster than another at a given temperature is said to be more reactive. Reactivity is measured by several methods, using carbon dioxide, air, or steam as reactants.9 ebastian and Mayers" have developed a method for the determination of absolute reaction rates between coke and oxygen by a study of ignition points under certain conditions. These and other investigators have established the relative reactivity of types of carbon. Lignite, charcoal, bituminous coal, cokes in the ascending order

Jan 1, 1951

By Y. Dardel

Since the first determination of Dumas1 in 1880, many authors have tried to measure the solubility of hydrogen in solid aluminum, or at least the amount of dissolved gas in it. However, the interpretation of the results which have been obtained is still difficult. Sieverts? then Baukloh and Oesterlen,a during their gassing experiments could not determine any solubility of hydrogen in solid aluminum. On the other hand, Portevin, Chaudron and Moreq4 leaving the determination of the hydrogen solubility beyond the scope of their experiments, measured only the amount of gas that they could extract by high voltage positive ion bombardment and found up to 151 cm3 of gas per I00 g of metal. Winterhager,6 by repeated vacuum extraction during severe cold rolling, extracted up to 161 cm3 per 100 g. In both series of experiments, the extracted gas contained chiefly hydrogen with a small amount of nitrogen, carbon monoxide, carbon dioxide and methan. Later, Schmidt and von Schweinitze as well as Chaudron and Moreau7 showed the importance of the adsorbed layer of hydrogen, the extracted quantities being proportional to the surface and not to the weight of the sample, when its cleaning was not perfect. In earlier experiments, Steinhausers extracted from more massive samples nothing but a little amount (< 0. 5 cm3 per I00 g) of hydrogen, carbon monoxide and methane. Thus, according to these experiments the quantity of dissolved gas is probably very small and the effect of the surface could falsify the results of the extraction. However, owing to the care taken by Chaudron and his coworkers, it was clear that all the hydrogen extracted could not come from the superficial layer, or that the links which bound hydrogen to the surface were so strong that the usual method of cleaning could not destroy them. After the main parts of the experiments reported here had already been carried out, Eborall and Ransley9 published the results which they obtained with an analysor having an accuracy of 0.03 cm3 per I00 g, while that of the usual apparatus was only about 0.2 cm3 per I00 g. During all their experiments the adsorbed layer would have been a difficulty, for the value of the adsorbed hydrogen was approximately equal to that of the dissolved amount. Table i—Amount of Adsorbed and Dissolved Hydrogen (After Eborall and Ransley) Extraction Hz Adsorbed „ h2 dissolved Temperature Calculated Extraction OCcma 500 0.057 0.050 600 0.062 o.096 Since always keeping the same adsorbed layer is very difficult, an error in the correction for the adsorption can be made when using the apparatus of. Eborall and Ransley, as well as the earlier apparatus.

Jan 1, 1949

By John W. Bell

IN the introduction to an excellent pamphlet, John Gross1 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 for 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.1 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 O. Gates2 and No. 23 is one by the writer.3 Employing the subject matter contained in papers Nos. 22 and 2323 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&apos;s was a. true hypothesis and Stadler&apos;s false. The evidence in a third paper4 will be employed to prove that Stadler&apos;s hypothesis, when examined experimentally, is worthless. STADLER&apos;S HYPOTHESIS WORTHLESS It is the writer&apos;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&apos;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&apos; Opinion What information is available in Mr. Gates&apos; 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&apos;s energy units are obtained per foot-pound-applied in coarse than in fine 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&apos;s theory is applied, the number of mesh-tons (my unit of surface produced) is

Jan 1, 1942

By Harold Vagtborg

Equilibrium vaporization ratios were obtained for light hydrocarbons, nitrogen, carbon dioxide, and hydrogen sulfide in a reservoir fluid containing 35 mole per cent hydrogen sulfide. The data cover the range from 700 to 2500 psia at 154°F. The results of this study are compared with results obtained by other investigators on systems containing less hydrogen sulfide. When large amounts of hydrogen sulfide are present, the equilibrium vaporization ratios for ethane and heavier hydrocarbons are greater than the ratios for systems containing little or no hydrogen sulfide. Large amounts of hydrogen sulfide have the reverse effect on the ratios for methane. Routine PVT studies were performed on the reservoir fluid including flash and differential liberations of the dissolved gas. The pressure-volume relationships of the saturated fluid were also determined. It is shown that the use of published correlations on physical properties of reservoir .fluids can lead to gross errors when applied to systems containing large amounts of hydrogen sulfide. INTRODUCTION Equilibrium vaporization ratios for various components in naturally occurring crude oil and condensate systems have been reported in the literature.1,2,3,4 Recently, several papers have been published with emphasis on non-hydrocarbon components such as nitro- gen, carbon dioxide and hydrogen sulfide.5,6,7,8,9 The increased discovery of reservoir fluids having high concentrations of these non-hydrocarbon components, particularly hydrogen sulfide, has created a need for a further study of their phase behavior. Equilibrium vaporization ratios are known to vary with pressure, temperature, and overall composition of the system.&apos;" It was anticipated that ratios for systems containing large amounts of hydrogen sulfide would differ from the ratios for systems containing small amounts of hydrogen sulfide. Therefore, the present work was undertaken in order to obtain reliable equilibrium vaporization ratios for the reservoir-fluid in question. PROCEDURE The procedure followed was evolved after considerable experimentation on methods of handling the very sour crude. It was found that drying agents (e.g. CaCl2 and CaSO4) absorbed an appreciable amount of hydrogen sulfide from the gas used to prepare the recombined samples. The concentration of hydrogen sulfide dropped as much as 10 mole per cent in a single pass through a 3-ft drying tube. It was also found that when undried gas was agitated with mercury in the PVT cell, a black scum, believed to be mercuric sulfide, was formed and the hydrogen sulfide content of the gas dropped several mole per cent. Because composition changes during a run could not be tolerated, separation of the mercury and sour fluid was necessary. In order to effect this separation, a stainless steel piston, equipped with Teflon O-rings, was placed in the PVT cell between the mercury and the sour fluid. This piston had to be constructed in three parts because

Jan 1, 1955

By A. Kenneth Schellinger, O. Cutler Shepard

Overgrinding, Or the breaking of ore particles into sizes smaller than required for liberation, is a first-magnitude problem in grinding for concentration processes. The conventional ball mill-classifier circuit used in commercial grinding produces mineral particles that are all ground to pass an upper size limit, while the proportion of very fine sizes, or slimes, is uncon rolled. Calculations based on measurements made in commercial plants which were grinding ore for flotation concentration indicate that about go pct of the energy used in grinding is consumed in useless overgrinding.1 In addition to accounting for the consumption of a large amount of energy, over-grinding is detrimental to the flotation process. Overground slime coats larger partitles and inhibits selective action by the flotation reagents on particles which by themselves could be separated very well. A further difficulty is that extremely finely ground minerals do not respond satisfactorily to flotation-collecting reagents, and both selection and recovery decrease markedly with particles finer than 5 to 10 microns.2 conventional classifier-ball mill circuits inherently tend to overgrind the valuable minerals more than the gangue minerals because most valuable minerals are of higher density and are softer and more friable than the gangue minerals. The rem- edy to this situation would seem to be a separating mechanism which would separate free valuable mineral particles regardless of size and density considerations and a mechanism which would operate in the grinding mill so that valuable mineral grains would be removed from the grinding machine as soon as they are freed. A consideration of these ideas led the authors to the conception of a simultaneous grinding and flotation mechanism. So far as the authors are aware, such a machine has never been tried, even on a laboratory scale. Development oF Apparatus and Experimental Method Apparatus development went through several stages before a really workable machine was devised. Work was started with a large cast-iron mortar 7½ in. in diameter and 8 in- deep. A cast-iron disk Was mounted on the end of a vertical shaft SO that it fitted loosely into the bottom of the mortar. On top of the disk four radial cleats ¼ in. high were bolted to agitate the ball charge as the disk, or impeller, rotated. The vertical shaft was driven at 100 rpm by a mechanism improvised from an old coffee-mill type of grinder. A hole was drilled into the bottom of the mortar through which compressed air was introduced for frothing. In operation, more charges of Ore which had been crushed to pass an 8-mesh screen were placed in the bowl along with water and suitable reagents. The impeller was then started and the compressed air turned On so that the ore was ground while being exposed to air bubbles. Mineralized froth was

Jan 1, 1949

By A. W. Schlechten, W. J. Kroll

As this paper is written, the only method for the commercial production of lithium metal is by the fusion electrolysis of LiCl-KCl mixtures, as first proposed by Guntz.2 The details of the industrial process have not been made public but Osborg8 stated that an efficiency of more than go pct is obtained with a lithium metal recovery above 95 pct and that the metal is 99.5 pct pure. Pletenev9 gave a power consumption of 66 kw-hr per kilogram of lithium metal produced with a salt consumption of 9 kg of LiCl and 0.4 kg of KCI. Pletenev&apos;s figures indicate a somewhat lower recovery than that given by Osborg. In a recent report on German lithium plants, Motock7 gives figures that are also less favorable than those reported by Osborg. At the German plants the lithium recovery in electrolysis was only 83.4 pct, with a power consumption of too kw-hr per kilogram of lithium. The metal averaged only 97 pct pure, with sodium and potassium as the chief impurities. The presence of SO2, SO3, Si02, Ba, Ca, and appreciable amounts of Na and Fe2O3 were especially disturbing in the electrolysis, and it was necessary to run the cell for some time to eliminate most of these impurities and condition the bath. It may be noted here that most of these impurities do not interfere with a thermal reduction process. These facts show that the present method of producing lithium is fairly efficient and would suggest that other methods could not compete. However, fusion electrolysis has certain drawbacks, such as the necessity of supplying costly, low-voltage direct current and of collecting and processing the anodic chlorine if it is not to be wasted. An even greater disadvantage is that lithium chloride must be used for an electrolyte. It is one of the most expensive salts of lithium, since it cannot be produced in anhydrous form by precipitation from aqueous solution, as is done with LiF or Li2CO3. The reduction of lithium by a vacuum process similar to the ferrosilicon method for the production of magnesium would have certain advantages. Small units, such as Pidgeon retorts, would give flexibility of production to meet demand, and a small plant would require moderate capital outlay. There is a possibility of using idle Government-owned magnesium plants for such purposes. The raw materials needed for vacuum thermal reduction would not have to be extremely pure, especially with respect to the iron and calcium content. This paper describes a series of experiments to determine the effectiveness of various agents for the reduction of lithium compounds. The investigation is part of a program of vacuum metallurgy con-

Jan 1, 1947

By George B. Clark

Many basic engineering principles of all four phases of mining operations, namely, prospecting, exploration, development, and exploitation, can be analyzed better in terms of quantitative geology. Geological data from both field and laboratory will also complement scientific methods now being developed. THE geological data emphasized so successfully in prospecting for new deposits, that is, structural controls, strength of solutions, and type of mineralization, are basically those required for successful exploitation. In the mining of newly discovered deposits the most economical methods should be employed as early as possible to keep the overall cost per unit produced at a minimum and to permit maximum extraction of valuable minerals. A crucial question is: How can geological data be translated into useful quantitative results which will aid in achieving this end? H. E. McKinistry&apos; has suggested that a solution may be reached in one of two ways: 1—the usual approach, use of judgment based on experience; or 2—mathematical calculations and tests on models, both subject to certain limitations. He also suggests that in addition to better use of geology more case data and theoretical data are needed on which to base sound judgment. Further research, therefore, is necessary. Perhaps in this field the emphasis should be on more specialization in mining methods and ground movement by men with thorough training in physics, engineering, geology, and underground mining. These specialists would be equipped to point out the most economical and scientific methods of exploitation. Selection of a stoping method is governed by the amount and type of support a deposit will require in the process of being mined, or by the possibility of employing the structure of the deposit to advantage in mining the ore by a caving method. In addition to these factors there are others which almost invariably influence the choice of an economical method of mining:&apos; 1—strength of ore and wall rocks; 2—shape, horizontal area, volume, and regularity of the boundaries of the orebody, and thickness, dip and/or pitch of the deposit and individual ore shoots; 3—grade, distribution of minerals, and continuity of the ore within the boundaries of the deposit; 4—depth below surface and nature of the capping or overburden: and 5—position of the de- posit relative to surface improvements, drainage, and other mine openings. In the final analysis it is usually necessary to disregard the less important of these factors to satisfy the requirements of the more important. Because of the variation of geological conditions throughout and surrounding the deposit, no mining method will be everywhere ideally applicable to the conditions encountered in one ore deposit. The immediate problem is to interpret the above physical characteristics of deposits in terms of geological characteristics. Very few quantitative geological data are available on the factors related to a choice of mining methods. However, there are many descriptive data in mining and geological literature which collectively show how important an effect details of geology have upon all phases of mining operations. The following categories of basic mining methods were investigated to establish the geological factors that have affected their successful application: 1— open stopes with pillars; 2—sublevel stoping; 3— shrinkage stoping; 4—cut-and-fill stoping; 5— square-set mining; 6—top slicing and sublevel caving; and 7—block caving. It should be noted that the first five of these methods are listed in the order of increasing support requirements. Mines were selected as examples only where geological descriptions were complete enough to warrant their use. A study of the geological factors involved in mining operations led to a choice of the following classifications, employed in Table I: 1—structural type of orebody; 2—dimensions (geometry); 3— country rock (type); 4—faulting, folding, and fracturing; 5—alteration of ore and rock; 6—type of mineralization; and 7—geological factors determining mining method (summary). Of these factors only one yielded results that can be defined from available data in a quantitative manner, i.e., dimensions of the deposit. These are the most reliable guides that can be used in selection of suitable mining methods. They are, in general, the properties of geologic structure most difficult to evaluate by studies of models, pho-toelastic studies, and other laboratory methods, all of which are at present more limited in their applications than the geologic method. Application of geology has proved a reliable guide in other phases

Jan 1, 1955

"The reduction works of the Boston & Montana Reduction department, near the north end of this dam is one of the reduction plants belonging to the Anaconda Copper Mining Company, the other being at Anaconda. These works employ about 1,100 men and have a capacity of about 3,800 tons of crude ore per day. This plant is now being remodeled to embody results of experiments and investigations which have been made with a view to more economic.al treatment of ores. A brief description of the plant as it will be after the remodelling is as follows:Concentrator Department: Consists of six units or sections, each section treating from 550 to 600 tons of crude ore per day. This department is not being. remodelled at the present time. A concentrating process, however, has been developed and partially installed in the old mill, and an experimental section embodying this concentration process has also been installed at the concentrator at Anaconda."

Jan 1, 1913