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By E. J. Wasp, N. T. Cowper, R. A. Davis, W. F. McDermott

On October 26, 1967, the world's first long distance iron ore slurry line was put into service by Savage River Mines. The commissioning of this $5 million installation climaxed three years of development, testing, and construction. The mine and concentrator are located approximately 30 miles from the town of Waratah in a rugged semi-mountainous terrain. The open pit mine will produce approximately 5 million long tons a year of crude ore, which is ground in autogenous mills, separated magnetically, thickened and prepared for the pipeline. The 53-mile pipeline transports the 2,250,000 tons of concentrate north to Port Latta. The Port Latta facilities include the receiving, storage and thickening equipment for the slurry as well as filtering, balling and pelletizing equipment to produce the 2,250,000 tons of high grade iron ore pellets. Five vertical shaft furnaces are used for pelletizing.

Jan 1, 1969

By Davis RA, Wasp EJ, Cowper NT

The transportation of solids by pipeline is not new. In 1891 a United States patent broadly covering a method of pumping coal with water was granted to W. C. Andrews. His pilot plant was constructed on what is now the site of the New- York Public Library in Manhattan. His process was applied to pumping coal mine waste back into the worked out portion of coal mines. On October 26, 1967, the world's first long distance iron ore slurry line was put into service by Savage River Mines. The commissioning of this 5.0 million dollar installation climaxed three years of development, testing, and construction and proved the Company's faith in pioneering' this unique transportation method for iron ore. This paper will discuss some of the more pertinent items in connection with that three-year period. Also to be discussed will be the operation of this line during the past eleven months.

Jan 1, 1969

By John B. Cook

Most of today's open pits take the form of conical-shaped excavations in the relatively flat or undulating terrain surrounding them. Ore is usually hauled uphill from the pit bottom by truck, rail, or skip to crushing, beneficiation, railroad loading, and stockpiling facilities near the edge of the pit. There are, however, a few orebodies whose locations in mountainous terrain causes special problems that generate different technical solutions. LAMCO's 10.5 million metric tons per year high-grade iron mine in the Nimba Range in tropical Liberia is probably one of the most extreme of these exceptions (Fig. 1). At Nimba there is a 2,000-ft difference in elevation between the upper- most levels of the orebodies and the start of the railroad below at the foot of the mountain. Haulage to the crusher is usually downhill at grades of up to 12% dictated by the topography of the rugged terrain. Ore is transported from the crusher down the mountainside by a conveyor sys- tem 10,096 ft long, whose average grade is 18%. The belt is braked by asynchronous generators, feeding power into the local network. The original crusher must be relocated at a lower elevation outside the final pit limit to allow that part of the mountain on which it now stands to be excavated. HISTORY The LAMCO Joint Venture is owned 75% by the Liberian-American- Swedish Minerals Co. (LAMCO) and 25% by Bethlehem Steel Corp. LAMCO is comprised of a 50% interest by the Liberian Government and 50% interest by Liberian Iron Ore Ltd (Fig. 2). This joint venture was launched to develop the high-grade iron ore deposits of the Nimba Range.

Jan 1, 1969

By John A. Reeves

The unit-train concept of moving raw materials .is considered by many railroad men as the most signifigant contribution to reducing rail freight costs since the introduction of the diesel locomotive. The idea, which involves assigning railroad equipment to only one haul and keeping the equipment rolling results in the maximum utilization of equipment. For example before the unit train was plaged in servige on coal hauled to one of Mid. Continent's customers, about 900 cars and 9 locomotives were needed to move the coal requirements for this plant, compared with. 250 cars and 6 locomotives used by the present scheduling of unit trains. If the economics of unit-train movements are to be achieved, coal companies no longer can use railroad cars as small storage bins on wheels to accumulate their coal in trainload shipments, but, rather, must build large bins which incorporate facilities for the rapid loading of railroad cars in a minimum length of time. In this manner railroad equipment is operated at a high degree of efficiency 'with consequent dollar savings. After negotiations between Mid-Continent, the customers and the Railroad, it was decided a unit train loadout and storage facility at Carbondale was required in order to reduce the delivered price of coal to the customers and thereby contribute to the reduction in the cost of manufacturing steel. Carbondale was selected as the bin site location, because the cost of constructing a railroad to Redstone could not be justified for the tonnages anticipated.

Jan 1, 1969

By Martti E. Maliniemi

The Svappavaara mine is located 47 km southeast of Kiruna C. The ore deposit now being worked, Leveäniemi, belongs to Luossavaara- Kiirunavaara AB (LKAB) and is the company's, and also Sweden's, third biggest mine after Kiirunavaara and Malmberget. The distance by rail to the Norwegian port of Narvik is 214 km and to Lulei on the Gulf of Bothnia 340 km (Fig. 1). The village of Svappavaara has about 800 inhabitants. The climate is arctic in character. The temperature during the months of December to February is about -20º to -35ºC, and the ground is covered by a blanket of snow for about eight months of the year. The brief, intensive summer often arrives with startling speed, bringing both sunshine and warmth. HISTORICAL BACKGROUND Mining operations took place in Svappavaara as far back as 300 years ago. During the period 1655-1684, copper ore was taken from the deposit called Gruvberget in the vicinity of the village. The total production of raw copper was probably about 1,045 tons. During the 18th and 19th centuries, several attempts were made to extract iron ore from the numerous deposits found within the district. These brave attempts, however, were foiled by the high phosphorus contents and the enormous transport difficulties. The present-day mine fields in Leveäniemi were discovered in 1897. At that time, it was difficult to determine the size and value of the deposit because of the thickness of the soil cover (3 to 15 m) and the

Jan 1, 1969

By R. S. Shoemaker, R. M. Darrah

Expanded markets for copper in the past few years and a consequent search for new ore bodies have revitalized the widely known but seldom applied method of producing copper called heap leaching. Heap leaching is defined here as the process applied to oxide ores which have been mined solely for the purpose of leaching. In this case, all costs of mining the ore become a part (and a substantial one, as will be shown) of the cost of the copper produced, whereas in dump leaching the mining cost has already been accounted for as a stripping cost for the sulfide ore. Of course, with newly opened ore bodies, stripping costs may be divided through accounting procedures, between the copper produced by the mill and that obtained by leaching, but in general this type of operation must be economic on the basis of the sulfide ore alone. Heap Leaching Requires Low Capitalization Heap leaching, as a beneficiation method, is unique in that it is the only process used for the recovery of a metal in which the time involved from mining the ore to metal production is measured in months instead of hours or at the most, days. Then too, after the ore body has been depleted and mining has ceased, metal production will continue for some substantial period of time. Heap leaching is also unique because unlike any other method of beneficiation, metal recovery cannot be accurately forecast ahead of time and recovery will never approach that of a conventionally milled sulfide ore or even that of a vat leached oxide ore. Counter-balancing this drawback, however, is the relatively low capital expenditure necessary to put a heap leaching operation into production and the consequence that smaller ore bodies which would not support the higher capital cost involved in a vat leaching plant can be made into producers. Stated more simply, in heap leaching the operator accepts low recovery in exchange for low capital costs and low break even point. Whether the operator chooses vat leaching with its higher recovery and higher capital costs, is a mater of economics which depends in part on the nature of the ore to decide. For instance, an ore that will vat leach in 7 to 10 days using 7 pounds of sulfuric acid per pound of copper produced might use double or triple that amount of acid during the months that are required for heap leaching.

Jan 12, 1968

By B. Mienie, J. F. Morrison, C. H. Wyndham, M. J. Martinson, N. B. Strydom

Contribution to discussion M. J. Martinson (Associate Member)*: Application of the simple selection test mentioned in the penultimate paragraph of the paper (reference to a more detailed description of the test would not have been inappropriate) should assist South African mining engineers to make more efficient and possibly more humane use of available labour. The authors mention that the test 'is in use in the gold mining industry', and it would be interesting to know the extent of this usage in terms either of number of mines or as a percentage of the total number of labourers employed on 'heavy' tasks. The paper would have been of greater interest had the authors attempted to relate their findings to the statistics of labour usage in gold mines, members of the Chamber, and to the 'heavy' tasks in particular. On this the authors confine themselves in their Introduction to the somewhat ambiguous statement that 'A relatively high proportion of the 360,000 Bantu mine labourers are employed underground on manual tasks'. To appreciate the practical significance of the authors' work one needs a detailed breakdown of the underground labour force by occupations, together with estimates of quantities such as: (a) total stope and development tonnages moved manually per shift: by shovelling by shovel/tram by tramming together with distances and any other relevant data; (b) theoretical energy expenditures required to accomplish (a); (c) actual energy expenditures required to a accomplish (a); (d) energy theoretically available to accomplish (a) based on 50 per cent of the total maximum oxygen intakes of the labour force allocated to these tasks. Analyses of this sort on an industry basis might demarcate the more promising areas for detailed investigation. Stress is often laid on the need for increased mechanisation to increase stope productivity, but it may be questioned if enough is known about the characteristics, performance and potential of our underground labour force to warrant blind acceptance of stope mechanization as the industry's panacea. Provided that health and safety are not compromised, the ultimate yardstick for assessing any form of stoping is profitability; obviously the industry must investigate very thoroughly any form of mechanization which might reasonably be expected to increase profitability, but it should also be remembered that a radical overhaul of existing systems-which have after all evolved by trial and error over the past 75

Jan 11, 1968

On the preceding pages, operations of the four major producers of nickel today have been described. There are, however, other operations in the world that, while not as large, are still of importance to the over-all picture of the industry. Among these are the Finnish producer, Outokumpu Oy, a state-controlled company that operates the Kotlahti mine 200 miles northeast of Helsinki. The ore from this operation, which began production in 1959, is pentlandite, chalcopyrite and pyrrhotite, occurring as disseminations in the host rock and in breccias. The flowsheet of the 1500-tpd mill at Kotlahti follows standard practices of crushing, grinding, and flotation to produce a pentlandite-pyrrhotite concentrate and a copper concentrate, both of which are shipped to the smelter at Harjavalta.

Jan 10, 1968

By J. H. French, O. E. Lissner

S. J. Venning (Member): The authors should be congratulated on producing such concise data from the plethora of experimental report results. Their observations on the cost of liners calls for some comment however, as the really unique feature about liners in South Africa is that they can be obtained so cheaply. Attention has been focussed on the 1961 cost of white iron at R50 per ton for chill cast block liners while manganese steel liners are reported as costing as much as R250 per ton. The former are however, a relatively heavy, simple casting with a low labour content and reasonably high percentage productivity, made from metal melted in a cupola. The capital investment to produce this type of casting is low compared with that required to produce manganese or other alloy steels. The chill cast liner block is always a production item and is cast in a permanent mould. Almost all manganese mill liners, except perhaps small aperture screens, are sold for considerably less than the R250 per ton quoted by the authors. Where a price of this nature is asked it would be because the design of the casting demanded is complicated or leaves little metal to carry the labour and other costs of production. As reference has also been made in the paper to liner costs in America, prices for similar types of liners obtained during a visit to Canada and the U.S.A. in 1961, could be of interest: Material Price per Location short ton Ni-hard R250 Elliot Lake Body liners block type Ni-Hard R300-310 FAS New York Mn Steel R280-290 FAS New York Screens (grids) slotted Cr.Mo. Steel R600 FAS New York hole Ni-Hard R320 Elliot Lake At this time the price of comparable manganese steel block type liners on the Witwatersrand was between R145/R155 per short ton. White iron in Canada was then quoted at about R150 per ton, but as Ni-Hard gave a life of at least two to twoand-a-half times that of white iron, it was a much more attractive proposition particularly when relining and down time costs were taken into consideration. There is little doubt that most foundries would welcome clear and complete specifications of the customers requirements and systematic checks to ensure that these specifications are being met. B.S.S. or other accepted standards should form the basis of these, and periodic inspections should be undertaken to ensure that only those foundries capable of complying should be included as suppliers. The authors have referred to new designs, but can the industry really call the grid and plate new when the patents were taken out during the second world war and the sawtooth block has been used for at least 35 years. Unfortunately, far too little thought and consideration are normally given to the ultimate mill liner design while the mills themselves are on the drawing boards. The position and size of the door opening, which is the key to the ultimate shell liner configuration, the drilling of the

Jan 9, 1968

The world's largest steam-driven mine hoist, built by Nordberg Manufacturing Co. a half century ago, has become a major tourist attraction in Michigan's Upper Peninsula. The hoist, designed by company founder Bruno V. Nordberg, is located in the Quincy Mine Number Two, just north of Hancock. Nordberg began construction of the mammoth machine before World War I, but delayed completion until after the war. The hoist became operational in 1920 and continued running until the mine was shut down in 1931.

Jan 9, 1968

By Plato Malozemoff

Our young, technically oriented people today are entranced by the space program, by physics that unlocks the secrets of nature, by electronics, and by other new technologies. The mining industry seems to be their last choice. Enrollment in mining schools is dropping and so is the interest of our great universities in mining curricula. The image of the mining industry appears to be one of being conservative, nonprogressive, resistant to change, and lacking in excitement and challenge. I submit that this public image is wrong. While the initial development of mines is usually slow and seemingly plodding, the problem of solving the imponderables of nature-the ore emplacement, its structure and physical characteristics, its response to recovery processes, and its probable ultimate profitability-is stimulating to the most acute and searching of minds and requires careful application of a combination of many technologies. The paths to obtaining the desired results must be within the bounds of reasonable expenditures in the early stage of development when risks are still high, and this financial discipline is a further stimulus to the imaginative search for the right answers.

Jan 6, 1968

By E. P. Pfleider, W. G. Pariseau

Even as the rational selection of excavation equipment requires a matching of machine performance capabilities to rock response characteristics, the functional features of transportation systems must be matched to the characteristics of the materials handled in order to secure optimum, trouble-free operation. In particular, ore passes and chutes must be designed for flow in order to insure controlled draw whenever a chute is opened. The alternative to quantitative design is, of course, the well known ad hoc procedure of freeing plugged ore passes, cleaning up spillage from flooded chutes, etc., after installation of the system. Since the voluminous empirical work sporadically undertaken through the years evidently had not led to reliable design methodology, it seemed worthwhile to attempt a rational and complete description of the movement of material through ore passes. The concepts of de- formable body mechanics applicable to gravity-induced flows of granular media or "soils" were involved. Accordingly, the mathematical theory of soil plasticity was adopted as a working hypothesis, and a two-dimensional laboratory model of an ore pass constructed in order to investigate the appropriateness of the theory. The general experimental strategy consisted of formulating the movement of material in ore passes as a boundary value problem in soil plasticity, determining the required boundary values of stress and velocity experimentally, and then using these data for the initiation of a theoretical solution for stress and velocity throughout the moving mass of material. A comparison between the theoretical and experimental constituted the test of the original hypothesis.

Jan 6, 1968

By Robert M. Grogan, James C. Bradbury

The Illinois-Kentucky mining district has, since 1880, accounted for 80 per cent of all U.S. production of fluorspar. The ore deposits are of two types: vein deposits formed by fissure fillings along faults and bedding-replacement deposits formed by the replacement of strata at certain favored stratigraphic positions along minor faults and fractures. Fluorite is the chief valuable mineral in the deposits. Sphalerite and galena commonly are present in small amounts but, in some ore bodies, may form a substantial part of the valuable constituents of the ore. Calcite is the chief gangue mineral; quartz and barite Exposed strata range from Devonian are present in various amounts. through Pennsylvanian in age. Igneous intrusive rocks occur as mafic dikes and sills and intrusive breccia dikes and plugs. Structurally, the district is situated on a generally northwest- trending arch that has been sliced into a series of long, narrow blocks by normal faults that trend northeastward. The structural The localization of the ore deposits by faults and fractures, the association of fluorite mineralization with intrusive breccias, and various lines of petrographic and geochemical evidence indicate that the fluorite-zinc-lead deposits were epigenetic and deposited by solutions emanating from a deep alkalic magma.

Jan 1, 1968

By Arnold S. Epstein

Tlze green electroluminescence occurring at liquid-nitvogen temperature in epitaxial gallium phosphide diodes is exarnined using the donor dopants silicon and sulfur. Zinc is used as the p-type diffusant. Two tt~ain peaks are found, namely til 0.56 and 0.535 p. The intensity of the gYeen peak occurring at 0.56 is found to oary as -4"' wlzere Jf is the forward current density and to first order appears to be dependent on the concet~tvation of ionized donovs. Tile vadintive lifetinze fov this process is -10-'sec. The intensity of the 0.56-p lu~ninescence is nt least several ovdevs of magnitude greate,, at liquid-nitvogen temperature than that noted jor the 0.538 radiation for current densities up to about 35 amp per sq cm. The variation oj- the intensity of. the 0.538-p peak with current density caries according to the donor dopant. With sulfur as the donor one has a J12 dependence, whereas when silicon is the donov a J1.5 dependence is noted. Tl~e above can be fitted to a In molecular process (in first order) such as a band to band recombination. The constant B fir band to band vecombination) is found from our experitt~etlt to be of the order of 2 CZI cm per sec, zcl~ereas for tlze impurity vecombination peak (0.56 p]) tile A cotzstor~t is of the order of. 7 x 10-' cu cm per sec. The structuve o/. the diodes at liquid-nitvog-en tett~peruture is oj tlze double injection type, and by carefi~lly selecting. the thickness of the structure one can regulate the onset of the 0.538-p radiation. The diode clzavaclevistics follow an I-Vx laiv where x is >1. Llregatiue resistance in these structures con be correlated with saturation of tlze 0.56- peak. ThE development of high-purity, large single crystals of epitaxially grown gallium phosphide is the ultimate aim of the studies to be reported. While a number of papers and reviews' have been presented on properties of gallium phosphide grown from gallium solution and from the melt, less attention has been paid to the properties of epitaxial gallium phosphide grown from the vapor. In the studies to be reported we have examined the effects of shallow donor impurities, primarily silicon and sulfur, at liquid-nitrogen temperature, on both the electrical and green electroluminescent properties of carefully characterized single-crystal epitaxial Gap. I) EXPERIMENTAL PROCEDURES A) Material Preparation and Properties. The samules used in this investigation were single-crystal n-type epitaxial gallium phosphide. he single crystals were synthesized from the elements by vapor transport and epitaxial deposition on single-crystal gallium arsenide substrates. Typical transport agents used were chlorides. The growth procedures are described in more detail in Ref. 2. The samples were grown in the (111) orientation and had thicknesses up to 480 . Following the growth, the GaAs substrate was removed from the epitaxial Gap by etching in nitric acid. The Hall effect and resistivity of these n-type samples were measured as a function of temperature and analyzed for impurity content and compared with spectroscopic emission analysis. The results are given in greater detail in Ref. 3 but a summary is noted in Table I. In Table I, AED denotes the donor ionization energy, Nd is the donor impurity concentration, lvA is the acceptor impurity concentration, NI is the total impurity concentration, Rip vs temperature is the measured Hall mobility-temperature relationship from liquid nitrogen to about 400"K, n(77"K) is the measured carrier concentration at liquid nitrogen, and R/p(77'~) is the measured Hall mobility at liquid nitrogen. Good agreement is noted between electrical and emission analysis for the major donor impurity, silicon, in the two materials where it is a major constituent. In the case of the two samples SC90-6 and AP138-3(111) B which were sulfur-doped the ionization energy found from the electrical data is in good agreement with that reported for sulfur.4 The acceptor impurities are not known, although from the presence of magnesium in the samples as indicated by emission analysis and its known role as an acceptor in GaP,4 magnesium could be present as an acceptor. Similarly, since silicon is known to be an amphoteric impurity4 it may also be present as an acceptor in the Gap.* The *A small band has been noted at 0.625 11 in samples SC90-6, SC76-4, and SC77-4, suggesting according to Ref. 5 that sillcon could be present as an acceptor In these samples since these authors have found this peak to represent a donor-silicon emission. presence of calcium deduced from emission analysis is not correlatable with the electrical analysis and its presence as an electrically active impurity should be questioned. B) Sample Preparation and Diode Fabrication. The original slices were of the order of 2 sq cm in area witi the (111)B face up. The material was cleaved and sectioned and the best pieces selected (i.e., those which visually appeared to have the least strain and defects). The samples were polished, lightly etched, and cleaned in trichloroethylene. Zinc was diffused in a closed tube at 800°C to form a p-n junction. Diodes were made from these samples. The depth of junction varied, but did not exceed 10 p depth. The back side of the slice was lapped back and a low-resistance n-type contact was formed (at 600°C in forming gas) to an Au-Mo tab with a hole in the middle for

Jan 1, 1968

By Donald F. Hammer, Donald W. Peterson

The Magma mine at Superior, Arizona, has produced over 13 million tons of ore yielding 1.5 billion pounds of copper. It is a mesathermal deposit, and, although the bulk of the ore has come from the Magma vein, most of the recent production has come from replacement deposits in limestone. The mine lies in a Precambrian and Paleozoic sequence of rocks that includes schist, diabase, quartzite, shale, and limestone; nearby, these rocks are intruded by igneous bodies of Mesozoic age that were probably the source of the ore-bearing fluids. Locally, these rocks are overlain by Cenozoic volcanic rocks and conglomerate. Layered rocks at the mine dip eastward about 30°. The area is extensively faulted, and surface faults are grouped into an earlier east-trending set and a later northtrending set. An additional set of diagonal faults that trends northwest and northeast is recognized underground. The east-trending Magma vein-fault, with its splits and branches, is the principal mineralized structure of the district. The bulk of the ore occurs in large ore shoots that plunge steeply westward and are sporadically connected by irregular ore shoots dipping gently eastward. The main ore controls apparently are selective zones of permeability related to the transverse faults. A bed of crystalline limestone near the base of the Martin Limestone (Devonian) is the host for the replacement ore bodies-discontinuous tabular pods dipping eastward parallel to the bedding. Selective permeability is again the apparent ore control, caused both by local hydrothermal solution of limestone and by deformation along faults. The principal copper minerals in the vein are chalcopyrite, bornite, enargite, tennantite, chalcocite, and digenite; these minerals vary in abundance according to a distinct zoning pattern. The upper part of the mine yielded considerable zinc as sphalerite, and the oxidized zone was rich in silver. The chief gangue minerals are pyrite, quartz, and hematite. In the limestone replacement bodies, chalcopyrite and bornite are the chief ore minerals.

Jan 1, 1968

By Gordon B. Brox, Joseph F. Mcaleer, Charles C. Goddard, Edward P. Shea, Robert G. Ingersoll, Lester G. Zeihen, George J. Burns, John M. Guilbert, Richard N. Miller, Charles Meyer

The Boulder batholith is a composite intrusive in which the Butte quartz monzonite is the dominant rock type. Quartz porphyry dikes intruded the quartz monzonite in directions which were subsequently followed by the eastwest- striking Anaconda system of veins, the earliest major veins in the district. Anaconda veins are faulted by northwest-striking veins of the Blue system, which are most prominent in the central third of the district. Anaconda veins are the major producers in the western third of the district and also in the eastern third, where the "horsetail" zones break off from them. Individual oreshoots are thousands of feet long as well as deep on both systems of veins, but they are generally more persistent and thicker on the Anaconda veins. Viewed broadly, the mineralization in the big veins is arranged in crudely concentric zones of zinc and manganese around a central zone of copper. Only slightly offset from these zones, but earlier structurally, is a zone in which quartz-molybdenite veinlets are prominent. This zone tops at about the 2800 level and widens downward. The pattern of distribution of Cu-Fe-S series minerals in the large veins of the copper zone is not symmetrical with respect to the metal zones. There is a fringe of chalcopyrite-bornite all around the copper front. However, within the copper zone, pyrite-chalcocite-enargitecovellite- digenite assemblages are prevalent near its eastern edge from the 3800 level to the surface, whereas deep on the large Anaconda veins in the western part of the copper zone, chalcopyrite is dominant. Here the chalcopyrite grades upward into bornite-chalcocite ores and eastward into the high-sulfur assemblages of the Leonard mine. Sericite-biotite-K-feldspar alteration envelopes, containing up to about 2 per cent copper as chalcopyrite, are present on some of the quartz-molybdenite veinlets at deep mine levels in the molybdenite zone. These envelopes contrast sharply mineralogically and chemically with the later sericitic and argillic alteration envelopes around Main Stage veins. K'0-Ar'0 ratios indicate that they may be older by about 4 m.y. Secondarily enriched copper ores are mined in the Berkeley pit in the southeastern part of the district. At the present time they account for about half of Butte's production of copper.

Jan 1, 1968

By R. T. Moolick, John E. Neill, O&apos

In selecting a particular stripping method the ultimate aim is the removal of material at the least possible cost. Accomplishment of this goal requires the careful consideration of many factors including geologic and topographic information, proximity of the stripping operation to waste disposal areas, possible problems of land reclamation, types of suitable equipment available, and auxiliary equipment requirements. 4.2-1. Selection of Stripping Methods. The size of an orebody and the distribution of values within that orebody normally will limit the variety of economical stripping methods which need to be considered. Much may depend upon the selectivity required in the mining due to the relationship between ore and overburden, as well as the character of the overburden itself. The following factors concerning the geologic nature and environment of an orebody, as well as the production requirements, must be determined before any selection of equipment is made: 1. The size of the orebody and distribution of the values within that orebody. Is the ore massive or scattered, bedded or disseminated, thick or thin? 2. The nature of the overburden to be removed. Is it a hard dense rock, bedded rock (thick or thin), friable material, earth, sand, elay, marsh, etc.? 3. The character and significance of geologic structures (fractures, faults, shear zones, etc.) associated with the ore occurrence. Are there water bearing formations with resulting water disposal problems? 4. In considering the nature of the overburden, including its alteration products, the physical or chemical conditions which, when combined with anticipated climatic conditions, may render certain equipment inoperable during unfavorable seasons.

Jan 1, 1968

By Kenneth K. Humphreys

INTRODUCTION What is automation? Why automate? Webster's New Collegiate Dictionary defines automation as "the automatically controlled operation of an apparatus, process or system by mechanical or electronic devices that take the place of human organs of observation, effort, and decision." This definition gives an implied answer to the second question, )"Why automate?" Automation, automatic control devices and process control equipment receive their justification in supplementing or replacing manpower, which results in decreased operating costs. In the past, processes were operated with a minimum of instruments for the purpose of merely guiding the operators. Processes were manually controlled by the operator on the basis of simple temperature, pressure, flow and level indications on a panel. Today the tendency is to design a new plant to operate in a fully automatic manner according to a preset pro- gram utilizing a minimum of manpower. In addition to this direct cost reduction savings in manual labor, increased equipment capacity results from uniformity of flow and uniform control of process variables. This control eliminates or minimizes surges and permits equipment to be operated at or near its ultimate capacity and, for a given material through- put, permits the use of smaller process equipment. Automatic control, therefore, be it in the field of coal processing or in any industrial field, is simply a problem of economics. Control equipment is justified in terms of its contribution toward increasing productivity, increasing profit margins or producing a new and saleable product at a profit. Instruments and control devices can be divided into two general groups: (1) units which measure, and (2) units which both measure and control a process variable. In coal processing, the first class of instruments, those which measure only, include analytical devices such as automatic ash, moisture and sulfur measuring units and the simple pressure, temperature, flow and level indicating devices mentioned previously. The second group of instruments, those which both measure and control, include a wide variety of units such as continuous weighing devices on conveyor belts and photoelectric cells which gauge the position of railroad cars on the track and insure that the cars are in the proper location for loading (see Chapter 15). The latter types of units as used in the coal industry differ from those used in other industries in one significant point. In coal processing these instruments need not be as highly sophisticated and accurate as in other industries; hence their cost is proportionately lower than in other industries. Because of the economic fact that coal sells for only a fraction of a cent per pound compared to dollars per pound for many chemical products, and

Jan 1, 1968

By G. A. Beske, P. Hilsch, J. Wulff

The effects of the growth precipitates on the su-perconduching properties of an Al- 15 al. pel Zu alloy have been studied using magneization, transition lem-perature, and residual resistivity measurements. Aging at 230C for 1/2 to 11 he produces an increase in Hfp, Ike field of first penetration, and tut increase in trapped flux. The upper critical field, Hc2 remains constant for such aging tunes. Aging at 200°C produces a decrease in Hfp and an increase in trapped flux. The transition temperature remains constant at Tc - 1.227° ± 0.02°K for aging at 220°C for tims of from 1/2 to 10 hr, and for aging at 200°C it remains constant at Tc. = 1.18° i 0.02°K for times from 1/2 hv to 26 hr. The observed behavior indicates that the superconducting matrix when aged at 220°C behaves tike a type U su-perconducior, but when aged at 200°C behares like a type I superconductor. The magnetization changes after aging can be attributed to the growth and clutnge in distribution of the precipitate in a matrix of fixed composition. IT has been realized for some time that the superconducting properties of alloys are highly structure-sensitive. This is evident in the work of Mould and Mapother1 who examined the properties of an aluminum alloy. Precipitation effects have also been briefly mentioned in other studies.2-4 Bonnin and coworkers5 have observed a sharp inc rease in trapped flux in an A1-Mg alloy when the residual resistivity reaches a value that indicates it is a type II superconductor. Recently Blanc, Goodman, and Nemoz5p have reported on precipitation effects in A1-Ag alloys. Detailed studies of precipitation in Pb-Sn and Pb-Cd alloys made by Livingston6 serve to show that the upper critical fields are enhanced by quenched-in solute. Furthermore. at various stages in the precipitation process? the upper critical field is determined largely by the remanent solute. The observed magnetic hysteresis and trapped flux in such alloys result from the interaction between precipitate and the flux filaments in the mixed state. The flux trapping appears to depend largely on the precipitate distribution and is greater if the superconductor is type II than if it is type I. To examine the effect of precipitation in another alloy system, particularly one in which the composition of the superconducting phase does not change significantly during the course of precipitation, the Al-Zn system seems interesting, especially in view of the transition temperature and resistivity studies of Chiou and seraphim.' Furthermore, the recent detailed electron-microscope studies of an Al-Zn alloy by Richards and Garwood8 provide a guide for relating superconducting properties to changes in structure. For these reasons, magnetization, transition temperature, and residual resistivity measurements for varying times of heat treatment at two temperatures were made with a -15 at. pct Al-Zn alloy. The results are reported in this paper. I) EXPERIMENTAL The alloy used was prepared by Cominco Products Inc. from 99.9999 A1 and 99.999 Zn. It was chill-cast, swaged, and drawn to 0.125-in.-diam rod. The analyzed composition was 15.6 at. pct Zn (31.1 wt pct). A jeweler's saw was used to cut the 0.125-in.-diam rod into the 0.875-in. lengths needed for magnetization measurements. The ends of the cylinders were slightly beveled to remove the sharp edges. A portion of the original 0.125-in.-diam rod was drawn into 0.010-in.-diam wire and used for transition temperature measurements. Specimens were suspended from wires within a vertical furnace during heat treatment. All were homogenized at 500°C for a minimum of 3 days and subsequently solutionized for times in excess of 6 hr. The solutionizing temperature was held constant to +3°C. Individual specimens were homogenized, quenched, and aged several times and no effects indicating loss of zinc were noted. The suspended specimens were quenched by cutting the wires and allowing the specimens to fall into a silicone oil bath held at the aging temperature. They were then isothermally transforked by aging in the same bath used for quenching. After aging for the required time (approximately 10 min aging was necessary before the upper critical field reached a fixed value), they were briefly (1 to 2 sec) dipped into an acetone bath at room temperature to remove adherent oil and then immediately transferred to liquid nitrogen where they remained until inserted into the helium bath (usually within an hour). The temperature of the aging bath was held constant to +3°C for longer aging times. The cryostat used for the measurement of superconducting properties consisted of outer and inner dewars for liquid nitrogen and liquid helium, Stokes 6-in. booster diffusion pump and Stokes Model 212 mechanical pump, temperature controller to measure and control the temperature of the helium bath, and a superconducting solenoid with inserted search coils to provide the magnetic field and measure the magnetic moment of the specimens. A schematic diagram of the cryostat is shown in Fig. 1. An ultimate temperature of 0.85°K was reached with this apparatus. The temperature controller used is similar to the one described by Blake and chase.' A 0.5 pct Allen-Bradley carbon resistor and a nichronle heater (200

Jan 1, 1968

By W. L. Sherwood

Present conventional techniques for primary steelmaking in both integrated and non-integrated plants are outlined. 'Direct reduction' alternatives are reviewed and categorized according to the nature of the product produced, with particular emphasis on the limited number designed to yield steel as the end product. A proposed new technique, referred to as the Sherwood Process, is described in some detail as a new combination of selected features from the prior technology. This process is in-tended to realize, in integrated operations, the result of continuons steelmaking directly from the ore in a single plant and, in non-integrated operations, continuons steel-making from a metallic charge of scrap, sponge iron or pig iron. Prominent features are continuons processing in elongated, horizontal, rotary furnace equipment, controlled direct heating of each process stage and, for the over-all integrated process, a size separation for removal of excess solid reductant and other waste just prior to the melting step. Material and energy balances are presented as well as exemplary cost data. Reasons for selecting this particular combination of process elements are dis-cussed.

Jan 1, 1968