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Greater automation of facilities is probably the best hope of the mining industry in its efforts to surmount continually rising labor costs. Underground mining, with its inherently low unit labor output, is especially vulnerable to rising costs, and here maximum automation is becoming increasingly important. One area of such operations that has been receiving a great deal of investigation is that of ore haulage and hoisting, two steps that often account for 10-35% of underground operating costs. Innovations and experimentation by mining firms and manufacturers over the past few years have resulted in larger, faster and more efficient tramming and hoisting equipment, but the White Pine Copper Co., at its operation in the Upper Peninsula of Michigan, is preparing to start up an entirely new concept in haulage later this summer. Known as the Dashaveyor System, it consists of individually powered modules (cars) that run on tracks inside a rectangular tube or on an open trackway at predetermined speeds up to 50 mph. At White Pine the Dashaveyor will be used to carry ore (crushed underground to -10 in.) from the mine to the concentrator. Dr. James Boyd, president of Copper Range Co., which is a 25% shareholder in The Dashaveyor Co., has indicated that if the Dashaveyor meets the expectations for surface operation, it will also be used underground for ore transport in place of the several miles of conveyor belts currently in use (Business Week, July 29, 1967).

Jan 7, 1968

By John V. Beall

What progress has been made in rapid excavation of tunnels and shafts? Where do we go from here? To get some answers to these questions was the reason that 500 tunnel men-contractors, geologists, professors, lawyers, mining engineers--came to the University of Minnesota May 15-17. The attendance of 500 astonished conference organizer Professor D. H. Yardley, who had anticipated half that number. They came from Japan, South Africa, Austria, Sweden, and Australia to name a few, and an officer of the Corps of Engineers thought all his people should have been there. The Civil and Mining Engineering Departments of the University combined for the program, which was staged by M. K. Cragun for the General Extension Division. The Saga of the Oso Tunnel The present state of the art of rapid excavation is "good." There is little controversy about where we go from here and what's more, this reporter finds that the technical know-how to automate excavation in any kind of ground is extant, even if it will take some years to achieve this state. In a nut- shell though, the present status can be seen in its various ramifications in the saga of the Oso Tunnel, as told by V. L. "Vic" Stevens of Boyles Brothers Drilling Co. :

Jan 7, 1968

By J. C. Draper

Duquesne Light Co.'s Portal No. 3 mine, in the Warwick group of mines in Southwestern Pennsylvania, was started in June 1965 with battery- powered tractor-trailer face haulage. The area as- signed to Portal No. 3 encompasses over 8000 acres of coal and will be developed by driving a set of main entries through the center of the tract in roughly a westerly direction, and sub mains on a 45º deflection angle in roughly a northerly direction (Figs 1 & 2). In the original planning these sub mains were to be at a 90º angle with the mains, but difficult roof conditions necessitated a change in plans. The entries and crosscuts are driven 18 ft wide in the normal Sewickley seam which varies between 46 and 56 in. high. The mine bottom is a strong shale and in average, cool or dry outside conditions stands up well for rubber-tired haulage equipment, but when the outside conditions are hot and humid, the bottom becomes soft and muddy. This condition lessens as the mine faces are advanced further from the portals because the condensation will then take place principally on the surfaces of intake entries outby the working areas. To aid the situation, an air cooling system was installed and is used as conditions dictate.

Jan 7, 1968

By N. E. Norman

During the past few years the underground mining industry and the big hole drilling industry have been involved in a flirtatious courtship, but until recently this courtship did not appear to be taken too seriously by either party. However, many of the promises made by the big hole industry are beginning to come true. Improvements have been made steadily in techniques of mechanical boring of large diameter shafts, raises and tunnels, thus making the big hole industry a more stable partner in the eyes of the mining industry. The mining industry too has made many changes in planning that make mining more attractive to the big hole industry. If there is to be a marriage between big hole drilling and mining, we should certainly take a look at the background of the big hole drilling industry, and examine the three main features of big hole drilling to determine how each applies to mining. Shaft Drilling Since 1950, numerous large diameter shafts have been drilled using rotary drilling techniques. Most of the larger diameter shafts have been drilled for the AEC for underground nuclear testing purposes. However, since 1960 several shafts have also been drilled for various mining companies, particularly in the United States. In 1967, rotary drilling shafts were either completed or in progress at the mining companies listed in Table I.

Jan 6, 1968

By James A. L. White

Of the natural elemental surfur deposits of the world, volcanic deposits are not only the most spectacular, but probably the most widely distributed. As defined for the purposes of this paper, volcanic sulfur deposits include those associated with all phases of volcanic activity, including eruptive flows, and fumarolic and hot spring deposits. The line of demarcation between the latter and sedimentary deposits is not always clearly defined. Although volcanic sulfur has played an important role for centuries, serving as essentially the only source of sulfur, currently it is considered as purely marginal material. Volcanic deposits were relegated to this position by the advent of Frasch sulfur into world markets, which established sulfur prices and annual production levels with which volcanic sulfur cannot compete, Since early in the 20th century commercial interest in volcanic deposits of sulfur has been high only when premium markets are available. This occurred in the early 1950s and is the case at present. Considerable effort is currently being expended to establish new producers of volcanic sulfur, and to expand current operations or reactivate previous ones.

Jan 6, 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 R. G. Beverly

The presence of radioactivity in uranium mill wastes has resulted in somewhat unique waste disposal methods. In addition to the common problems of disposing of large quantities of solid wastes, neutralizing acids, minimizing dissolved heavy metals, and clarifying all liquid effluents, the uranium mill operator must sample and analyze liquid effluents for micro-micro-quantities of radio- nuclides such as Ra226, Th230, and Pb210. Special disposal methods or decontamination procedures are required to meet the stringent limitations established by the U.S. Atomic Energy Commission on the concentration of radionuclides in liquid effluents. Routine monitoring of the receiving stream is also normally required. In recent years there has been a growing concern about the radioactivity remaining in uranium mill tailings, particularly at the sites of shut-down or dismantled mills. The State of Colorado, in cooperation with the industry, has adopted regulations for stabilization of inactive uranium mill tailings.

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 S. Terry, J. G. Munro

A unique feature of the recently dedicated pelletizing complex of Hamersley Iron Pty. Ltd. is the dry grinding circuit for the iron ore fines. This facility, located at Dampier, some 800 miles north of Perth, Western Australia, has a rated annual capacity of 2 million tpy of iron ore pellets. The Mt. Tom Price orebody, which supplies the pelletizing plant, is a massive deposit of high-grade hematite containing an estimated 500 million tons of ore averaging 64% iron. The main impurities are ferruginous shales that occur in bands of varying widths in the orebody. An average analysis shows 4.4% SiO2, 2.0% A12O3, 0.05% P and 1.8% LOI.

Jan 5, 1968

By E. Wroth, D. G. Krige

Written contribution E. Wroth* (Visitor): The technical aspects of the assistance scheme which IS provided for in the Gold Mines Assistance Act, 1968, and certain consequential amendments to the Income Tax Act, 1962, are explained by Dr Krige. The object of this contribution is to explain the working of the Gold Mines Assistance Act (hereafter referred to as the 'Assistance Act') for the benefit of those concerned with the tax and assistance assessments of gold mining companies. The views expressed do not necessarily reflect the Revenue Department's interpretation of the legal provisions involved. 1. MINES WHICH QUALIFY FOR ASSISTANCE 1.1 The assistance scheme is applied on a selective basis. Mines which qualify for assistance are defined in the Assistance Act as 'assisted gold mines'. The requirements for classification as an assisted gold mine are that: (i) the mine be a producing gold mine; (ii) it would be likely to close down within eight years if not assisted; and (iii) with State assistance its life would be appreciably prolonged and there would be a significant increase in production of gold or uranium, or of both gold and uranium. 1.2 A mine is only considered for classification as an assisted gold mine on application. Such applications must be made to the Mining Leases Board through the Government Mining Engineer. Classification of a mine as an assisted gold mine is at the discretion of the Minister of Mines who acts in consultation with the Minister of Finance and has regard to the recommendations of the Mining Leases Board. 1.3 The Minister of Mines will impose conditions subject to which a mine is to be classified as an assisted gold mine. The most important will be that the mine must immediately lower its operating pay limit, as normally calculated, by a percentage fixed by the Mining Leases Board and conduct its future mining operations in accordance with the average grade of its ore reserves determined on the basis of such lower pay limit. l.4 The Minister will fix the date as from which classification as an assisted gold mine is to be effective. So long as the mine has operated on the basis of the lowered pay limit for at least three months during the particular year of assessment, he would normally fix the effective date as the first day of the year of assessment.

Jan 5, 1968

By G. A. Wiebols, N. G. W. Cook, N. C. Joughin

Discussion R. E. Rarnes (Member): The original concepts and the pioneering work now brought to the practical test stage by the Mining Research Laboratory team deserve our highest praise. The authors of this comprehensive paper rightly stress the urgency of establishing the extent to which the apparent potential can profitably be realised in practice. It is to be hoped that adequate funds will be made available by individual mining companies, the Chamber of Mines and manufacturers to attract the necessary staff and maintain the high rate of achievement of the last two years. From the Seventy-Eighth Annual Report of the Chamber of Mines and its members we see that, in 1967, with a Working Revenue of R759.8 million from gold and R54.6 million pit mouth coal sales and with profits from gold and uranium and pyrite of R307.9 million, only R1.9 million was spent by the Chamber on all forms of Research. It is considered unlikely that the associated mining companies and manufacturers exceeded this investment expenditure. Assuming a total of R4 million spent by the industry on Research and Development, this is less than half of 1 per cent of sales of gold and coal. This percentage, so low in comparison to North America and Europe, is no worse than that of Exploration expenditure which, in 1967, with South Africa's total mineral production of R1,287 million, was estimated to have been R6 million (Pretorius 1968). In a primary industry with ever present depletion of ore deposits and with cost escalation, expenditure on Exploration and on Research and Development is not a risky luxury but a tactical obligation. The potential rate of return on research expenditure into rock breaking is high. Stores consumed by gold and coal mines, members of the Chamber of Mines, totalled R316.2 million in 1967. Except for purchased power costing R42.2 million the highest cost group was explosives, drills and drill steel totalling R33.8 million or 10.7 per cent of the total stores consumed. In the paper under discussion it is claimed that the low 'effective stoping width' should greatly reduce the likelihood of rock falls or rock bursts to the extent that permanent support can be dispensed with. Insofar as this narrow cut is only 12 in. in advance of the working area which, with a 10 in. channel, is unlikely to be much reduced in width from that achieved by current methods, this claim is not readily understandable. Were it to have been based upon the regional support gained from packed waste it would have been more acceptable. Pre-developed stope drives may give serious trouble at depth and for this reason it is questioned whether a stoping area can avoid periodic sub-development blasting-the spoil and fumes from which will interfere with the rock flow and continuous mining of the rock-cutter. If, in the mining method proposed by the authors, stope drives are cut as small as possible (6 ft by 6 ft) then 25 per cent of the total tons handled (excluding resued waste) and 5 per cent of the gold will be blasted conventionally in the stope. At this stage one cannot envisage tunnelling machines economically or practically capable of such work. The various methods described by the authors and subsequent contributors for breaking waste are most interesting. It was noted that the 'bull wedge' and 'explosives' in Fig. 1 of the paper were no further from the ideal point 'A' than was 'cutting'. The writer considers that the bold and imaginative steps taken by the Mining Research Laboratory Team, the mining companies and the manufacturers concerned will eventually lead to a successful rock cutting machine with universal application largely independent of rock type. This may take many years. In the meantime other methods of improving productivity of saleable metal by rock breaking teams should be investigated even if such methods have local applications only. In 1955 the writer conducted tests with a wire saw similar to those used in quarries in the Northern Transvaal and elsewhere. Jeppestown shale, the immediate footwall of much of the East Rand gold field, was cut at the rate of 6 in. per hour using sand, water and a special endless rope driven by a low h.p. motor. A hypothesis on its application was submitted to the Office of the Government Mining Engineer in 1955 and to other mining institutions in 1966 after the writer returned to South Africa. By inference, rope sawing was classed as less promising than other methods tested in the Orange Free State Goldfields (Parker 1969). With highly resilicified hanging and footwall quartzites this was not surprising and confirmed the writers findings when testing hanging wall quartzite from the East Rand in 1955. The relatively uniform conditions, the low strength, hardness, and silica content of the Merensky Reef platinum deposits (Gray and von Bardeleben 1969) and in particular, the existence of overlying Merensky pyroxenite (Cousins 1964) make this and the East Rand attractive areas for larger scale testing of wire saws. It is envisaged that in suitable rock types the 5/8 in. slot would be advanced down dip or down a minor dip. In undisturbed areas 'faces' of up to 200 ft in length could be cut several feet in advance of breaking which could then consist of light blasting to the second free face or some of the methods now being tested for breaking waste in rock-cutting operations. A wire saw is an inexpensive and simple machine which, in some areas, could make significant and early gains in rock breaking efficiency as well as in ground and stoping width control.

Jan 5, 1968

By Edward A. Vitunac

The emerging importance of pneumatic conveying systems for bulk materials is largely due to the economic advantages such systems frequently exhibit when compared to alternate methods. Included in these advantages are relative simplicity, ease of installation, flexibility, ease of maintenance, cleanliness, safety and ease of automation to the degree desired. In general, installed costs favor a pneumatic system when compared to competing mechanical systems if the transporting distances are long or to and from many points. In addition, supplemental use such as cooling, heating, or drying may supply the incremental value required to justify its selection.1 The main disadvantages of a pneumatic system are the greater horsepower required per unit weight of material conveyed, material limits in abrasiveness, and density. In addition, material breakage tends to increase at changes of direction and at any obstruction such as that caused by the imperfect butting of two pipe sections in the conveying line. Classification of the material at the discharge may also cause problems in dusting and segregation. Materials weighing less than 4 lb per cu ft tend to cause feeding and collecting problems although they can be conveyed easily. Material between 4 lb and 20 lb per cu ft can be handled with fewer problems, but flow rate is hard to predict. The 20-60 lb per cu ft range offers best results, whereas materials of over 60 Ib per cu ft tend to drop out of the air stream and settle in the pipeline unless careful pro- visions are made to prevent this.2

Jan 5, 1968

By G. L. Wilhelm

Prior to the construction of a 450 tpd concentrator at the Mayflower mine, owned by the New Park Mining Co. and operated by the Hecla Mining Co., the common mining methods were cut and fill, with waste or gob filling and square set timbering with waste fill. Stoping operations were beset with the usual problems related to waste filling methods- storage and handling of fill material from development, costly waste excavations necessary to obtain a portion of the needed waste rock, slow mining cycles due to filling and inefficient ground support. After Hecla Mining Co.'s successful utilization of sandfill in its mines in the Coeur d'Alene district, utilization of hydraulic fill in stoping operations at the Mayflower was a logical development of the modernization plan then in progress. The mine, located in the southeasterly part of the Park City mining district, Utah, is a medium size producer of lead and zinc ores, with gold, silver and copper as secondary minerals. The mine has been producing commercial ore since 1939 from a dual vein system which outcrops at an elevation of 7200 ft in the Wasatch Range. Access to the present mining areas is gained by a 7200-ft haulage tunnel and by an interior, three-compartment main shaft. The present mining horizons are about 2400 ft deep. The average horizontal cut and fill stopes, ranging from 3-15 ft wide and up to 120 ft long, can be classed as small.

Jan 5, 1968

By D. C. Seidel, E. F. Fitzhugh

Nickel sulfide ores lend themselves readily to concentration before smelting or pressure leaching, but neither of the major oxidized ore types-the iron-rich laterites and magnesium-rich, soft silicates-has been directly concentrated. Upgrading of crude ore in current practice is limited to cobbing lumps of the harder, lower grade rock (protore) from the softer, enriched silicate ore, because only a small percentage of the nickel in these ores occurs in discrete, contrasting mineral particles. There is no apparent mineralogical contrast in the superficial, iron-rich lateritic ores and, in fact, there is no firm assurance that the nickel atoms are within the lattices of the limonite minerals, or else irregularly adsorbed on the limonite. Within underlying, "rotten rock" silicates, occasional veinlets of garnierite and related nickeliferous silicates are found.1 The garnierite veinlets are only incidental, and the majority of the nickel atoms occur as erratic replacements of magnesium atoms in the micaceous chlorites of the ore mass.

Jan 4, 1968

By John D. Erickson

The concepts of long-range, open pit planning presented here are not new. The new idea is how the factors that control the location of the final pit limits can be considered altogether as one large, simultaneous equation. The X, or unknown quantity that is being solved for, is a long-range pit design that fulfills the desired objectives. The objectives of a final pit plan are determination of the extent and location of ore reserves, the allowable stripping necessary to uncover the ore, the final pit surface intercept, final pit topography and the contained metal values in both ore and waste.

Jan 4, 1968

By E. V. Howard

During the recent expansion of precipitate copper production at the Chino Mines Division of Kennecott Copper Corp., Santa Rita, N. Mex., local studies of the dump leaching process were intensified. As these studies progressed, it became apparent that a reliable method for locating leach solutions within the dump was required, since the contact between the leach solutions and the copper-bearing ma.teria1 determines recovery. In the past, the problem of determining leach solution contact has been attacked by various methods, including destructive testing (dump dissection), indirect observation (such as chloride or dye tracers) and mathematical or laboratory model studies, but they have not proved completely satisfactory. The investigation of methods for in situ observation of leach solutions led to consideration of radiation logging in drillholes, which are subsurface techniques used in petroleum exploration and development for locating fluids and measuring porosity. A significant advantage of radiation logging is that, unlike other types of logging, measurements may be made in cased holes, since gamma rays and neutrons will penetrate steel and cement. Three techniques are currently used at Chino: Natural gamma ray logging; neutron-neutron, or moisture logging; and gamma-gamma, or bulk density logging.

Jan 4, 1968

By L. B. Beneitone

Over the years, the rope changing procedure on the two double-drum San Manuel ore hoists has been streamlined and improved, so that a job that once required eight hours to complete is now finished in four. The ropes, constructed of improved plow steel, 6 x 25 filler wire, pre-formed, right lay, Lang lay with a fiber core and weighing 8 ½ lb per ft, have a 2 ¼ in. diam and are 3200 ft long. They have widely varying lives, but on the average a rope change on one of the two hoists is required every six months. When a rope change becomes necessary, the ropes on both drums of a hoist are replaced. Thorough planning and preparation are essential for efficient rope changes. A briefing session is held on the day prior to the change, during which the entire procedure is reviewed on a blackboard and discussed among the cable rigging crew. These discussions have led to many improvements that have been incorporated into the present rope changing procedure.

Jan 4, 1968

By Francis Cameron

Mining is both exciting and rewarding-although at times somewhat frustrating-and we all can have real pride in our industry, in its people, and in its accomplishments. It is, however, with concern that I have noted a deterioration in this Country in what might be called mining's stature and in the growth of a belief in many quarters that our mineral reserves are rapidly approaching exhaustion. In other words, there is a popular image that we are fast becoming a "have not" nation in many respects and that the domestic mining industry can no longer be considered, in the vernacular of Wall Street, to offer much in the way of "growth potential." I do not subscribe to this hypothesis, nor do I believe that the record of the mining industry bears this out. However, let me add that we can, in time, talk ourselves into this frame of mind and we can hasten the day when this very well might come about by unnecessary and unwise legislation or regulation. My remarks today are basically designed to give my reasons for refuting this negative philosophy and to review our record. With your help, I know we can improve our image, and the public's recognition of our industry's peculiar problems.

Jan 4, 1968

The new, internationally-owned Savage River Mines iron ore mining, concentrating and pelletizing project, built and operated by Pickands Mather & Co. International, was dedicated March 6 at Port Latta, Tasmania, Australia. Construction of the $79 million project began in 1965 and involved an open pit mine and concentrating plant plus a complete town for about 350 employees at Savage River; a 53-mile pipeline to transport the concentrates; a pelletizing plant and shipping facilities at Port Latta, which is on Bass Strait separating Tasmania from the Australian mainland. The project has a rated capacity of 2.25 million Ltpy of high-grade magnetite pellets, which will be sold to five Japanese steel companies under 20-year contracts. The open pit mine is located in the central portion of the Savage River deposits. Mining equipment, capable of moving 10 million tpy of ore and overburden, consists of four 6-cu yd electric shovels and 12 50-ton trucks. The concentrating plant, with a design capacity of 5 million tpy of ore, will use magnetic separators to upgrade the ore, which averages 38% iron, to a concentrate averaging 67% iron.

Jan 4, 1968

By E. Alan Haley

Accurate and economic control of line and grade on long tunnels, large structures, excavations and waterways has been a problem to surveyors and engineers for decades. The conventional method of surveying, incorporating a level, transit and a number of surveying technicians, has proven to be a very expensive and time-consuming operation. Now, however, engineers have adapted the laser to help ease the problems of surveying, at the same time increasing the efficiency of such operations. The laser, which projects a continuous, narrow red beam of brilliant light, provides a constant alignment reference that does not interfere with personnel or equipment. A complete laser alignment system, specifically designed for and suited to a variety of mining and tunneling operations, has been developed by the Perkin-Elmer Corp. The engineering significance of this development can be fully realized when it is noted that, aside from potentially tremendous economic advantages in terms of labor savings, the laser beam provides a system of alignment that is at least ten times more accurate than presently used optical systems. The new laser system provides accuracies within 10 microinches per foot (1 x 10-5in.) when used with an accessory centering detector. In one of its first field trials, the laser was used for alignment of a two-mile tunnel which showed a total deviation of less than 5/8 in.

Jan 4, 1968