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By M. T. Simnad, C. E. Birchenall, M. H. Davies

THE marker movements observed by Davies, Simnad, and Birchenall during the growth of magnetite on wustite have been misinterpreted. It is the purpose of this note to correct the original interpretation. In that investigation wustite specimens were prepared by oxidizing pure iron disks in water vapor until the iron was consumed. Argon was used as a carrier gas for the water vapor and on the complete conversion of the metal to its oxides ('FeO' plus a small amount (10 pct) of surface Fe3O) the water supply was bypassed and the specimens were annealed in argon alone in order to equilibrate the wustite with the overlying magnetite. After cooling, the magnetite was ground off, a crude magnetic test being employed to determine its complete removal. Radioactive silver streaks were then applied to the surface and the specimen was oxidized completely in water vapor to prevent hematite formation. After complete oxidation the markers were found to enclose about 80 pct of the magnetite with the other 20 pct lying outside the markers. This observation was interpreted to mean that oxygen transport was four times as great as iron transport during the growth of the magnetite. The conclusion would have been correct if, as in the case of wustite growing on iron, all the iron had started from one side of the marker and all the oxygen from the other side initially. Actually a material balance is required. For simplicity, take wustite to be exactly FeO, and magnetite to be Fe3O,. Then at the start of the reaction 3 FeO + O -> Fe,O., All of the iron and three-fourths of the oxygen of the final product will be within the marker inter- faces. The other fourth of the oxygen is in the gas phase outside the marker. From the experimentally determined final position of the markers simple calculation shows that one-fifth of the total iron has crossed the marker interface, while one-twentieth of the total oxygen has crossed in the reverse direction. The simplified result favors iron transport in the ratio four to one. However wiistite is not stoichiometric FeO. At 900°C, the temperature of the measurement, the composition is about FeO When the material balance for the same experimental result is repeated, assuming the product will be 1 mol of Fe3O4, it is found that 0.6 gram atoms of iron and 0.19 gram atoms of oxygen must have passed the marker in the same direction. (This is readily apparent from the reaction to form 1 mol of Fe3O4 SFeO The total oxygen supplied from the atmosphere is 0.61 gram atoms, but the iron which passes the marker interface, i.e., 0.6 gram atoms, requires 0.8 gram atoms of oxygen to convert it to Fe,O,.) If all the transport during magnetite growth is due to iron ion diffusion, then the marker would enclose about 85 pct of the final magnetite and the original measurement is in error by about 5 pct. This error is probably less than the uncertainty in locating the markers by autoradiography. The possibility also exists that the starting wustite was not quite saturated with oxygen and underwent further enrichment with consequent envelopment of the marker before magnetite caught up with the marker interface. In any case it appears that these measurements actually favor iron ion diffusion in magnetite overwhelmingly as the factor controlling its growth, although further experimental measurements are needed to eliminate the error in material balance described. Such studies are planned.

Jan 1, 1954

By D. Turnbull, J. C. Fisher

ACCORDING to the "reaction-path" modell,2 of martensite nucleation, the shear angle of the embryonic martensite plate must be treated as a variable, and included in any calculation of nucleus critical size. Also, as can be deduced from this model, the interfacial free energy between austenite and martensite does not reach its final value until the shear is completed. It is zero for zero shear angle. However, in order to account for the kinetics of the martensite transformation, some sort of interfacial energy barrier appears to be necessary even with the reaction-path model, for otherwise the volume and the energy of formation of the critical size nucleus both collapse to zero.3 Cohen independently suggested that surface energy could be incorporated into the reaction-path model, with the overall free energy of a martensite embryo being a function of its volume and shear angle.' It is possible to estimate the energy associated with the formation of a critical-size martensite nucleus starting with the reaction-path model and including a surface free-energy barrier. As the dependence of interfacial free energy upon shear angle is unknown, a simple type of dependence will be assumed, with the belief that the true dependence would not lead to appreciably different results. Consider the work required to form a lenticular martensite plate with radius r, thickness t, and shear angle 8. There are three contributions; one being the interfacial free energy, one being the free energy change in the martensite plate, and one being the free energy increase in the surrounding austenite. The interfacial free energy u is assumed to depend upon the shear angle 0 according to the relationship s=s0(?/?0)n [1] where 8, is the equilibrium shear angle and n is an exponent that may lie in the range 0 n 2. The work required to form the interfaces of a martensite plate then is W. = 2pr² s0(?/?0)n [2] The free energy change per unit volume of martensite is composed of two parts, one the ordinary volume free energy ?f1. which is negative, and the other the elastic strain energy G?m²/2, where G is the shear modulus and 7, the shear strain relative to the martensite structure. This expression for the strain energy is valid only when the shear strain ym, is sufficiently small that the martensite is within its linear elastic range. There is no doubt that ym, lies beyond the linear elastic range for embryos that are considerably subcritical. However, for critical nuclei it will be shown that ym, is 1.5 pct or less, within the linear elastic range of martensite. For embryos of nearly critical size, then, the strain energy of the martensite is correctly given by G?m²/2. The shear strain in the martensite is ym, = 8, — 8, and the work required to form the strained martensite is Wm --= (pr²t/2) [?fv + G(?O - ?)²/2] [3] The free energy change in the austenite is entirely that due to elastic distortion. The elastic strain is not uniformly distributed in the austenite, being large near the martensite plate and small elsewhere. Approximately, however, the energy corresponds to a uniform shear strain ya= (?t/2)/r [4] throughout the volume 4pr³/3 surrounding the plate. The work required to strain the surrounding austenite then is Wa = (4pr³/3) (G?a²/2) = (G?²/6) prt² [51 For simplicity, the same shear modulus G is assumed for each structure. The total free energy for forming a plate then is W = W3 + Wm + Wa. = 2pr² s0 (?/p?0)n + (pr²t/2) [?fr+G(?0-?)²/2] + (G?²6) prt2 [6] This expression is correct for nuclei and for embryos of nearly critical size, where, as will be shown, the strain energy in the martensite is correctly given by the expression G (? — ?)². Having W as a function of r, t, and 8, as in Eq. 6, there is a saddle-point where W has a stationary value, W subsequently decreasing indefinitely as the nucleus volume increases along the reaction path. The stationary value of W is the energy of the critical nucleus. The critical nucleus has radius, thickness, and shear angle such that ?W/?r - awlat: = ?W/?p? = 0. Performing these differentiations and calculating the critical nucleus energy, W* = [8192p(G?/6)²;s/27 ?fv4] [7] where a= (?/?0)3n+1[l +G(8"-8)'/2af.]' [7a] and where 8 is to be determined from the equation (1 + 3n/4) + G8(6O - (9)/[Af. +G(6>o-6>)72] = 0 [8] For ?f, near —200 cal per mol or —10" ergs per cc, and 8, near 1/6, as for iron-base alloys, Eq. 8 gives ?0 - ? ~ - (4 + 3n) ?f1./4G0O [9] as the difference between the equilibrium shear angle and the actual shear angle for a critical nu-

Jan 1, 1954

By A. J. Barry

An apparatus and procedure have been developed for testing the anchorage effectiveness of slotted-type bolts. The apparatus consists of: (1) A hydrualic Jack assembly for loading the bolt and (2) a device (extensometer or cathetometer) for measuring displacement of the bolt as it is loaded. This apparatus and procedure permit measurment of anchor slip for loads up to the yield load of the bolt. Two representative makes of slotted-type bolts (a total of 94 bolts) were tested in typical sandstone and shale mine roof. Analysis of the test results showed that (1) anchorage capacities up to the yield load of the bolts can be developed in either rock type; (2) the shape of the slotted end of the bolt affects the ease with which the bolt can be driven; (3) all bolts for which the driving distance was more than 1.25 inches produced satisfactory anchorage in both rock types (thus, driving distance can be used as a criterion for satisfactory anchorage); (4) for the group of bolts that was satisfactorily anchored the anchor slip was unaffected by the type of bolt but was considerably larger in shale than in sandstone; and (5) the bolt expansion, that is, the amount the wedge expands the prongs in excess of the diameter of the hole, can also be employed as a criterion for satisfactory anchorage. Although the bolt-expansion and driving-distance results do not differ appreciably, the latter is recommended because it is easier to determine.

Jan 1, 1954

By B. M. Hamilton

"IntroductionTHE PROCEDURE and technique for drilling holes has probably undergone as great a change during the past ten years as any one of the other operations conducted in hard-rock mining. During this period, we have seen a change from the heavy leyners and drifters to the light jackleg type of drill machine; from steels that were forged and hardened on the bit end to the inserted •carbide-tipped rod, from the necessity of providing special bit-end fabrication and beat treatment of shank steels to the simple grinding or turning of a plain section on a shank with no heat treatment required for bit attachment.When the steel maker was first called upon to manufacture a hollow drill rod, it was required that the steel must not only have sufficient strength and fatigue resistance to cope with the stresses of service, but also must be capable of being heat treated on the bit end to a high hardness for resistance to wear and abrasion . Little was known at that time about alloy steels and the beneficial effects to be gained from additions of chromium, nickel , and molybdenum. So it was that the plain 0.80 per cent carbon analysis was selected, since it had already proven to be a suitable grade for chisels, hammers, and other such shock tools. This grade has long endured the test of time and is still used in considerable tonnage today, particularly by the construction industry.As the metallurgist learned more about the influence of alloying elements, a new hollow drill analysis was placed on the market in the early 'thirties'."

Jan 1, 1954

By H. A. Johansen

Recent advances in the metallurgy of chromium have enabled this metal to begin to take its place among the refractory metals. Considerable work is being done on the development of high-temperature alloys containing high percentages of chromium.1/2/ In spite of the relatively large amount of work done on these alloys, little is known of the properties of the pure metal. Information in the literature concerning the tensile properties of pure chromium metal is quite scant. This deficiency maybe explained by the great difficulty of preparing test specimens from such a brittle material, resulting in many losses in each phase of the specimen preparation and testing procedure.

Jan 1, 1954

By L. H. Van Loon, E. G. Crayston, A. F. Heather

"IntroductionTHE PROPERTY of Madsen Red Lake Gold Mines. Limited. comprises fifty-eight claims located between Russet and Faulkenham lakes in Baird and Heyson townships of the Patricia portion of the District of Kenora. The mine is seven miles •by road from the hamlet of Red Lake. This latter centre is .connected to the Trans-Canada Highway at Vermillion Bay, Ontario, by provincial highway No. 105 which has a total length of one hundred and thirteen miles.The Madsen property went into production on August 11th, 1938, with a milling plant rated at 300 tons per twenty-four hours. Changes made in the crushing plant during October, 1938, allowed a •finer feed to be delivered to the mill and increased the milling •capacity to 400 tons per twenty-four hours. During 1948, steps were taken to add an additional 400-ton capacity unit to the milling plant. This latter work was completed by May, 1949, giving the present total milling .capacity of 800 tons per twenty-four hours.It was realized in the early 1940's that, eventually, considerable material suitable for backfilling in the mine would be required. As there were no large deposits of gravel or sand anywhere near the mine, officials of the Company considered it advisable to investigate the feasibility of using mill tailings for mine •backfill. Since Madsen mill tailings contain approximately 80 per •cent of minus-200-mesh material, it was not considered practical to use them for underground backfill without first removing the finer material. as the slimes present in a product of this fineness would without doubt preclude any drainage whatsoever. Table I gives a sizing analysis of Madsen mill tailings and Table II a quantitative chemical analysis of the ore."

Jan 1, 1954

By H. M. Thorne

The Union of South Africa has been interested in oil shale for a long time, and deposits of oil shales, torbanites, and similar rocks have been found in many parts of the country. (Fig. 1.) Samples differ markedly in oil yield and in the characteristics of the oil. The strata vary in thickness, and most of the deposits that have been found are in unfavorable locations because of remoteness from markets or absence of transportation facilities. This report deals mainly with the torbanite deposit near Ermelo, Transvaal, in the Union of South Africa, and the operations of the South African Torbanite Mining and Refining Co., Ltd. (hereafter referred to as Satmar), in mining the torbanite, retorting it, and refining the crude oil.

Jan 1, 1954

By G. Vibert Douglas

"Gold occurs in Nova Scotia in three ways, saddle reefs, fissure veins, and fossil placers. Of these the saddle reefs are the commonest. They have been the source of most of the gold and have also been the subject of most of the literature on the gold occurrences.The Meguma or Gold-bearing series of Nova Scotia is a conformable succession of quartzite's and slates, possibly Precambrian in age, which have been folded into a number of anticlines and synclines that strike (Fig. 1) roughly east and west. This folding has been termed Acadian and is probably a westward continuation of the Caledonian revolution in Britain. It may have started in the Ordovician and continued through the Silurian and into the Lower Devonian. It was followed by the intrusion of the great Devonian granite batholiths. It is thought that the mineralization is connected with the late phases of the cooling of the granite. As the name saddle reef implies, the veins form on the crests and flanks of the anticlines, possibly in the synclinal troughs as well. The beds favoured by the solutions are the slates. The reason for this is apparent when it is realized that in order to fold over 30,000 feet of sediments (the thickness of the Meguma according to E. R. Faribault) through p radians of arc it is necessary to have a differential movement between the top and bottom beds of p times the thickness of the sediments. In the case of the Gold-bearing series, in which strong quartzite beds lie between incompetent slate beds, most of this differential movement was taken up in the slates. The evidence for this is the remarkably fine drag folding and the intense shearing and comminution of some of the beds of slate (Fig. 1). The quartzites exhibit some shearing and faulting, but there is little doubt that the greater permeability lay along the incompetent slate beds. The mineralizing solutions followed these beds and also passed from one slate bed to another through a joint or fault in the intervening bed of quartzite. These joining veins are called angulars in Nova Scotia. Within the veins the gold is in shoots and the real problem is to find out what controls the position and extent of the shoots."

Jan 1, 1954

By N. J. Grant, D. S. Bloom

N recent reports, Sully' and Beck and coworkers' I have advanced hypotheses concerning the formation of the phase. Both of these hypotheses are based on Pauling's theories of the electronic configuration of the elements of the first transition group. Sully considered that the number of electrons which can be absorbed in filling electron "holes" is the factor which determines whether or not the structure can form. Beck suggested that the presence of a certain concentration of electron "holes" is the controlling value. It can be shown that a similar criterion can be developed without making reference to Pauling's theories. Table I shows the elements of the first transition group and the incidence of the a structure in the binary systems of these elements according to the data presently available. Also appended is the number of electrons considered to exist in the 3d-4s levels of these elements. If the atomic percentages for the phase are listed as shown in Table 11, and if the number of electrons in the 3d-4s levels per atom is calculated for the a boundary values, the results are as shown in the last column. It can be seen that the numerical values are all in the vicinity of 7 electrons per atom. This is true also for the ternary Cr-Mo-Ni IT phase, which occurs in a system in which none of the binary compositions are known to develop the phase. In the Fe-Mo binary, the high temperature a phase, which occurs at a composition of 50 atomic pct of each element, also yields a ,;1 e of 7 electrons per atom. From inspection of Tables I and 11, it develops that any series of sequential numbers which is assigned to the elements under consideration will result in the numbers pertaining to a compositions grouping themselves around one specific number. It does not follow, however, that this random number necessarily has physical significance. Such a point has already been discussed by Hume-Rothery." One interesting point develops from this study, however, and that is the position of manganese with reference to the other elements and to the phase. In the system used here, the formation of phase is associated approximately with the value of 7 electrons per atom, which is also the number of electrons per atom that manganese has. Manganese, it will be recalled, exists in three different crystal forms, none of which are the usual, simpler metallic forms; on the other hand, elements just to the left of manganese (in Table I) generally have lower coordination-number structures (body-centered cubic) than do elements to the right of manganese (face-centered cubic). The contribution to a complex crystal structure by the interalloying of elements from opposite sides of manganese is significant in that the a phase is formed in binary systems including one element from each side of manganese, with manganese capable of participating with both sides. It is indeed an interesting coincidence that manganese, with 7 electrons per atom, solidifies in complex crystal forms, and that when alloys of the transition elements are made which have seven 3d-4s electrons per atom, they, too, tend to assume a complex crystal form. From the above, it might be deduced that the study of the a phase might be advanced by a critical examination of the element manganese and how its structure is affected by small additions of elements which change the 3d-4s electrons per atom number. The role of atom size would also be of interest in formation. References ' A. H. Sully: The Sigma Phase in Binary Alloys of the Transition Elements. Journal Inst. Metals (1951) 80, Part 4, p. 173. 2 S. Rideout, W. D. Manly, E. L. Kamen, B. S. Lement, and P. A. Beck: Intermediate Phases in Ternary Alloy Systems of Transition Elements. Trans. AIME (1951) 191, p. 872; Journal of Metals (October 1951). "W. Hume-Rothery, H. M. Irving, and R. J. P. Williams: The Valencies of the Transition Elements in the Metallic State. Proc. Royal Soc. (September 1951) 208A. p. 431.

Jan 1, 1954

By Arthur B. Yates

"The nickel deposits of Sudbury, discovered in .1885, have probably commanded as much attention as any single group of orebodies in the world. They have been the major source of the world's nickel supply since 1903, when their production first surpassed that from the Pacific Island of New Caledonia. From 1903 to 1914 the production from New Caledonia remained essentially constant, whereas that from Sudbury increased from 13,000,000 to 50,000,000 pounds a year. After 1914 the production from New Caledonia gradually dropped to about 4,000,000 pounds in 1919, where it remained until 1923. During these years the Canadian production increased steadily and since 1923 has gone up at an amazing rate to a peak of over 280,000,000 pounds in 1941. Meanwhile the New Caledonia production increased to about 10,000,000 pounds until the pressure of war raised it to roughly 14,000,000 pounds a year.The Sudbury deposits, because they have been the source of 80 per cent of the world supply of nickel, because of their economic importance to the world of metals, and because they are geologically unique in their occurrence, structure, and genesis, have attracted foremost consideration and the literature relative to their geology is voluminous. This is particularly true in regard to genesis.There is a significant and interesting parallel in time between the advancement of the science of economic geology and the great expansion of the nickel industry. The controversies regarding the fundamental classification of orebodies into generalized types were in full swing during the height of the development and expansion of the great nickel mines, and the Sudbury deposits, being the most important and the only ones of the sulphide type extensively developed, have been the chief source of information in regard to this type. The observations and the conclusions drawn thereon from numerous studies of the various mines of the district have not only influenced, but are the essence of, this chapter of the science."

Jan 1, 1954

By J. J. Forbes

This publication is a revision of the first of a series of four Miners' Circulars originally prepared and issued in 19'29 for use in a course of training designed to prepare mine officials to organize men for rescue and recovery operations. These circulars were revised in 1937 and 1938 and again in 1948. This circular (Miners' Circular 33) discusses mine gases and methods for detecting them. The second of this series (Miners' Circular 34) gives instruction in Bureau of Mines methods of sampling and analysis of mine gases; the third (Miners' Circular 35) describes methods and equipment used for protection against mine gases; and the fourth (Miners' Circular 36) explains procedures used in sealing and unsealing mine fires and in recovery operations after mine fires or mine explosions.

Jan 1, 1954

By Gerald C. Mathis

FERGUS County, Mont., shown in Fig. 1, is known for its once famous gold mines near the old towns of Gilt Edge, Maiden, and Kendall. But at Heath, a small farming community near the foothills of the Snowy Mountains, thousands of tons of gypsum have been mined since 1916. The value of the gold taken from the mines of Fergus County runs into millions of dollars; the value of the products made from the gypsum of the Heath mine is also in the millions. The gypsum mine will continue to produce for many years, while the gold mines are closed today. The Heath gypsum mine was developed by the Northwest Gypsum Co. in 1916. The products of the small mill built by this company were plaster, agricultural gypsum, and raw gypsum for cement manufacturing. The United States Gypsum Co. prospected the adjacent property in 1916 and 1917 by means of two adits and some exploration drill holes. During World War I, the company abandoned operations at Heath, but again became active in this area in 1928 hen it purchased all the property of the Northwest Gypsum Co. A long range program of expansion was begun with the building of a block plant and the purchase of additional land and mineral rights. A modern wall board plant was built in 1936. When the old wooden mill purchased from the Northwest Gypsum Co. burned in 1937, an all-metal calcining plant located at Blue Rapids, Kansas, was dismantled, moved to Heath, and erected on the site of the mill. In August 1951 a new mine was opened 3000 ft north of the mill. Rock is conveyed from the mine to the mill by an 18-in. belt conveyor. The Heath gypsum deposit lies at the base of the Ellis formation of Jurassic Age. The Ellis at Heath lies unconformably on the black fissle shale of the Heath formation of the Big Snowy Group, which is Mississippian in age. The strike of the gypsum bed is N45°E and the dip approximately 3" NW; the regional strike and dip of the area. The gypsum lies in a series of six layers separated from each other by thin seams of shale. The lower layer is composed of' gypsum and shale intermingled and is not of commercial grade. The upper five layers are high in purity, but only the middle four layers are mined on the advance, as the top layer is left for roof support. This top layer is mined on retreating operations at the same time the pillars are robbed. Total thickness of the gypsum layers varies from 12 to 14 ft. Fig. 2 shows the average thickness and arrangement of the layers. Mining Methods The mine was originally developed by driving an adit into the gypsum horizon. The Northwest Gypsum Co. mined in the top four layers, the top layer being split, leaving about 12 in. for roof support. A modified room and pillar method was used: entries were 15 ft wide and rooms varied between 20 and 30 ft and were turned on 90" from the entries. Crosscuts between the rooms left pillars from 10 to 20 ft square. The rock was loaded by hand into 1-ton mine cars and pulled to a central gathering point by mules; from the gathering point, mules or horses were used to pull the string of cars to the mill tipple. This method was continued by the United States Gypsum Co. from 1928 to 1937.

Jan 1, 1954

By T. R. A. Gokcen

THE possibilities of separating and purifying metals by high vacuum distillation were examined by Kroll.1 He suggested vacuum treatment for the removal of zinc from the lead produced after Parkes desilverizing. The St. Joseph Lead Co. developed the first commercial vacuum dezincing process at their Herculaneum refinery, as described by Isbell.² The Broken Hill Associated Smelters at Port Pirie, Australia, has, after several years of pilot plant operations and fundamental investigations, developed a continuous process, which will be described briefly in a forthcoming publication." As the full-scale continuous vacuum dezincing plant at Port Pirie is still experimental, publication of full practical details of the plant will be deferred until the unit is operating as a normal part of the continuous refinery. This paper deals only with theoretical aspects of vacuum distillation processes, with particular reference to vacuum dezincing. The method of mathematical analysis is of general interest as it may be applicable to other metallurgical separations which have been investigated recently.4-6 Evaporation Processes At about atmospheric pressure, or higher, most liquids possess a boiling point—a temperature at which any heat put into the liquid is absorbed only as latent heat, not as specific heat. If a steady heat input is supplied, the liquid's temperature rises to this value, then remains constant while bubbles of vapor form beneath the surface. The rate of evaporation is determined solely by the rate of heat transfer to the liquid; the temperature of boiling is determined by the partial pressures of the volatile constituents in the liquid, and the total pressure above the surface. If the rate of heat transfer to the liquid is increased, the temperature remains constant, and the rate of boiling increases. When evaporating metals under vacuum, however, the partial pressures concerned are generally so small that boiling does not occur, because at even a fraction of a millimeter below the surface the hydrostatic pressure is usually too great to permit the formation of a pocket of vapor. In addition, the high thermal conductivity of metals tends to prevent the local superheating which is necessary for bubble formation.' Although this effect is doubtless also exerted when boiling metals at higher pressures, the magnitude will be less because the degree of superheat required to form bubbles is very much less at the higher temperatures involved. Under vacuum, therefore, evaporation of volatile constituents takes place only from the exposed surface, and the rate of evaporation depends upon the surface area, the surface concentration of volatile constituents, the surface temperature, and the partial pressures of volatile constituents immediately above the surface. If the heat input is raised above a certain level, the effect is not to increase the evaporation rate at constant temperature, but to raise the temperature of the liquid until at some higher level an increased rate of evaporation (and thus of latent heat absorption) again balances the heat input rate.? Many substances (including metals) have a very large intrinsic evaporation rate at quite low temperatures—far below their normal boiling points. However, at atmospheric pressure, the large numbers of atoms evaporated are almost completely deflected back into the liquid (or solid) surface by air molecules. Thus the back condensation rate is practically equal to the gross evaporation rate, and the net evaporation rate is practically zero. It can therefore be seen that an overall distillation rate depends not only upon the intrinsic evaporation rate, but also upon the ability of the volatile atoms to move away from the evaporating surface. This movement is facilitated by the provision of a condensing surface close by. Vacuum Distillation: The function of the vacuum above the evaporating surface is to remove foreign molecules, so that the chances of deflection of an evaporated atom back into its source are reduced. When the residual gas pressure is reduced so far that the evaporated atoms have a high probability of reaching a nearby condensing surface without suffering collision with a foreign molecule, the state of affairs is termed "molecular distillation." This process is practiced commercially today for the purification of numerous organic chemical products of high unit value, but not, to the writer's knowledge, for any metallurgical separation. When the degree of vacuum produced in a still is not sufficient to promote molecular distillation, then the evaporated molecules must diffuse through the

Jan 1, 1954

By W. J. Wrazej

THE measurements on the formation of multi-layered scales on pure iron1 and the metal -lographic examination of the produced layer of FeO, Fe,O,, and Fe2O3 shown in micrographs2 throw much interesting light on this apparently well-known subject. Professor Paidassi announced2 that in a few months7 time further results are going to be published with regard to the growth of Fe2O3 and Fe2O3 layers on an Armco iron as a function of temperature. Because of the very thin layers of Fe2O3 which are shown by Paidassi (Fig. 10 of ref. 2) when samples were scaled in air at 700°C for 4 hr and because this is not in agreement with the present author's previous observations, the following experiment was recently carried out. Polished cylindrical samples 13/64 in. in diam and 1/4 in. long from electrolytic iron used for spectro-graphic work (Hilger, H.S. Iron rods, F.693) and an Armco iron sample 1/4 in. in diam of the same length were suspended in a muffle furnace on Nichrome wire and annealed in air for 4 hr at 700°C 5OC. Because of the known easy peeling off of the Fe2O3 layer which is mainly in the form of powder (purple red dust on the surface) the samples were protected against any draft and shock. After annealing, the samples were cooled in air and placed in the mount-

Jan 1, 1954

By W. L. Bancroft, Peter Price

"What is now known as the 'Waite section' of the property of Waite Amulet Mines, Limited, was originally the Waite-Ackerman-Montgomery mine. It includes a group of twenty-eight claims lying on either side of the Duprat-Dufresnoy township line, and about 8 miles north-northwest of Noranda, Quebec. ,The original discovery of ore on the property was made by Thomas Montgomery in Block Q, Duprat township, and was staked by him on behalf of himself, I. C. Waite, and C. H. Ackerman, in the summer of 1925. The property was optioned late in 1925 to N. A. Timmins who, after trenching and drilling the deposit, sold his 85 per cent interest to Noranda Mines, Limited. Early in 1927, a company known as Waite-Ackerman-Montgomery Mines, Limited, was formed and shaft sinking and development undertaken. Production commenced in August, 1928, ·and continued tJ May, 1930, when, owing to falling metal prices, shipment of direct smelting ore was stopped. Development of the mine by shaft sinking, drifting, and drilling continued until March, 1932, when owing to the prolonged low price of metals, the property was closed down. In September, 1933, an amalgamation with Amulet Mines, Limited, was effected, the new company being called Waite Amulet Mines, Limited. Early in 1937 operations were resumed, first by direct shipments to the smelter and later by aerial tramway to the Amulet mill. Operations have been continuous up to the present time, reaching an average of 720 tons a day milled for the year 1940. Production during the initial period (1928-1930) amounted to 75,811 tons, having an average grade of 7 .11 per cent copper, 4.47 per cent zinc, and 0.03 oz. gold and 0.65 oz. silver per ton. During the second period (1937 to end of 1946) , 1,060,524 tons, having an average grade of 4.53 per cent copper, 2.97 per cent zinc, and 0.03 oz. gold and 0.65 oz. silver per ton, were produced."

Jan 1, 1954

By R. H. Walpole, J. M. Kane

IN all probability the mining and metallurgical industry as a whole can demonstrate a larger ecorlomic return from installation of dust-control equipment than any other major industrial group. This fact has partially accounted for the marked increase of dust-control installations made during the past decade. While the primary objectives for installation of dust-collecting systems are improved working and operating conditions for men and equipment, the fact that an economic return can be anticipated on salvageable materials is an added advantage which shows in partial or complete equipment write-off. The conditions apply to most phases of the mining, milling, and smelting industry, both non-metallic and metallic. As with any mechanical devices, selection of suitable dust collector equipment involves evaluation of available products with characteristics most nearly meeting conditions of the application at hand. When there is valuable product to be collected, and/or when there are possibilities of air pollution or public nuisance, collector selection is often guided by the maxim of "highest available collection efficiency at reasonable cost and reasonable maintenance." A brief review of dust collector designs will permit outlining of major characteristics of each group. Final selection will involve detailed data against a background of the problem under consideration. The dry centrifugal collectors, see Fig. 1, represent a group of low cost units with minimum maintenance. They are subject to abrasion under heavy abrasive dust loads and to plugging with moist materials. Efficiency drops off rapidly on particle sizes below the 10 to 20 micron group. Because of the large amounts of —10 micron particles in most mining dust problems, they will normally be used as primary collectors and will be followed by high efficiency units. This combination is cspecially popular where the bulk of material is desired in a dry state with wet collection indicated for the final cleanup portion. In remote plant locations, dry centrifugal~ can be used alone if product in dust form has no value or if dust loading is light enough to eliminate a nuisance in the plant area. Where high efficiency dust colleotion equipment must be selected, choice will normally involve fabric arresters, wet collectors, or high voltage Electro-Static precip-itators. Fabric arresters, see Fig. 2, rely on the passing of dust-laden air at low velocity through filter fabric. Velocity ranges from 1 to 3 fpm for the usual installation and may be as high as 10 to 20 fpm in arrangements where automatic frequent vibration or continuous cleaning of the filter media is employed. Fabric is normally suspended in either stocking type or in an enlvelope shape. Collection efficiency is excellent even on sub-micron particle sizes. Equipment is bulky, must be vibrated to remove the collected dust load, and is restricted in applications from temperature and moisture standpoints. Condensation of moisture on the fabric filter mcdia causes plugging of the passages with great reduction in air flow. Temperatures for the usual medias of cotton or wool are 180" and 200°F maximum, although the introduction of synthetic materials such as nylon, orlon, and glass cloth have increased the possibilities of this type of collector for higher temperature applications. The wet-type collector may employ a number of different principles so that entering dust particles in the gas stream are wetted and removed. Principles usually include impingement on collector surface or water droplets, often in combination with centrifugal forces. Variety of wet collector designs is indicated by typical collectors illustrated in Figs. 3 and 4. Collection efficiency is a function of the particular design, although the better collectors will have high collection efficiency on particles in the 1-micron range. Wet collectors have the advantage of handling hot or moist gases, take up small space, and eliminate secondary dust problems during the disposal of the material. At times collection of the material wet is a disadvantage. Wet collectors may also be subject to corrosion and freezing factors. The high voltage Electro-Static precipitator, see Fig. 5, is probably the most expensive type of high efficiency collector. It finds its applications generally in problems in which collectors previously discussed cannot be employed. Its collection efficiency is based on its design features and can be excellent on the finest of fume particles. Material is normally collected dry. Gas temperatures are of no great concern as long as condensation does not occur within the dry type of precipitator and the temperatures do not exceed the limits for materials used in its construction. As with the fabric arrester, provisions

Jan 1, 1954

By Fred H. Klaer

Water supplies dependent upon induced infiltration can be developed by vertical wells, by infiltration galleries, or by horizontal water collectors. Each method is described and the advantages of water obtained by induced infiltration with respect to cost, chemical quality, and temperature are shown by specific water supplies. THE subject of ground water in industry is assuming increasing importance as new processes and new industries are developed. Industrial specifications of chemical quality and temperature for process water and for cooling purposes are becoming more rigid, and water must now have many properties to meet these requirements. As ground water is usually more constant in quality and temperature, many industries are turning to ground-water supplies to satisfy their growing need. As many published articles have stated during the past several years, the ever increasing demands have caused shortages of water supply in many localities throughout the United States. Although such shortages are generally of a local nature, they have given rise to many attendant problems, such as increased pumping costs, salt water encroachment, and rising ground-water temperatures. One partial solution to the problems of overdevelopment of the ground-water reservoirs in many areas can be accomplished by developing ground-water supplies that are dependent upon induced infiltration and by planning the use of underground reservoirs. These reservoirs can be filled during periods of excess precipitation and stream flow and can be emptied during periods of drought, in a manner similar to the use of surface reservoirs. In this way water that otherwise would be wasted into the oceans is salvaged during periods of excess. The development of water supplies dependent on induced infiltration is a relatively new concept and should be defined. When a well is pumped, the water level around the well is lowered in the shape of an inverted cone or cone of depression. As pumping is continued, the cone of depression will continue to expand, horizontally and vertically, until the quantity of water crossing the perimeter of the cone plus the recharge within the area of the cone is equal to the quantity of water being pumped.

Jan 1, 1954

By H. W. St. Clair

This study was undertaken to provide a more adequate knowledge of the chemistry of the sulfide volatilization process for recovery of tin from low-grade Bolivian ores. The investigation was partly supported by the Department of State under its program for scientific cooperation with South American republics. The results of this investigation demonstrate the conditions under which cassiterite (Sn02) may be converted to stannous sulfide by reaction with pyrite or sulfur vapor, Quantitative data are given on the rate of sulfidization and on the vapor pressure of stannous sulfide.

Jan 1, 1954

By R. C. McMurchy

"The Powell mine is in western Quebec about 2 miles northwest of the adjoining towns of Noranda and Rouyn. It is the original discovery of the area, its gold-bearing quartz vein having been discovered in 1922. Considerable work was done on this vein by several companies and discontinued because the grade, when gold was $20 an ounce, was too low for profitable extraction. Between 1922 and 1924 an inclined shaft was sunk to 225 feet and drifts were driven 240 feet on the 100-foot level.In 1929, due to the discovery and development of the Horne mine at Noranda, interest was revived in the property. Sulphide deposits of the disseminated and vein types were explored on the Powell claims. On the B copper 'showing', a vertical shaft, now known as the 'old copper shaft', was sunk to ·a depth of 167 feet and 300 feet of lateral work was carried out on the 150-foot level. As no orebody was discovered this work was discontinued.In 1933 the present company, Powell Rouyn Gold Mines, Limited, was formed and the gold-bearing quartz veins were again tested by drilling and trenching. In 1936, the price of gold having risen to $35 an ounce, operations were resumed. Since then the inclined shaft (No. 1) has been deepened to 500 feet, a second shaft (No. 2, vertical) on the main quartz vein has been sunk to 2,630 feet, and ll; shallow vertical shaft (No. 3) on the same vein but in the south claim has been sunk to 500 feet.Lateral work now approximates 9 miles, and to November 15th, 1946, 1,685,674 tons of ore, which averaged 0.129 oz. gold per ton and 73.6 per cent. Si02, had been extracted."

Jan 1, 1954

The Black Hills Mineral Atlas briefly summarizes all pertinent date available on the mines and mineral deposits in the Black Hills that have been explored, developed, mined, or located. The purpose of the atlas is threefold: First, from available date the general potentialities of the Black Hills region can be determined, although it is not possible to calculate actual ore reserves; second, it will provide the necessary basic information regarding a given property or commodity for further detailed study; and third, the compilation is a comprehensive directory of Black Hills deposits.

Jan 1, 1954