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By A. J. Richardson

Before the development of the underground stoping and mining costs can be considered, certain facts about the ore body, the proposed mine, markets, etc., must be known or determined. In the case to be studied, the zinc-lead mineralization occurred with a narrow vertically dipping structure of undetermined length and vertical extent. Exploration completed to date has revealed 6.5 mil¬lion st t of proven reserves. A further 820,000 st of in¬dicated reserves has been outlined and this tonnage is considered capable of being expanded by a factor of approximately four after more detailed drilling. After studying the market conditions and completing a very preliminary feasibility study, it was decided that production would be 730,000 stpy (or 2000 stpd) of ore. First year production would be at the rate of 1500 stpd. The main design criteria for the selection of the min¬ing methods are minimizing surface subsidence, maxi¬mum recovery of the ore body, maximum degree of grade control, maximum productivity, and safe working conditions. Two basic extraction systems are considered capable of meeting these requirements: mechanized cut¬and-fill stoping and sublevel long-hole stoping with filling. The primary development system of the mine has been designed to give maximum flexibility in stoping systems and layout and to permit changes if considered necessary as a consequence of actual production ex¬perience. At the present time, access to the mine is by a circu¬lar concrete lined vertical shaft, 16 ft diam, sunk to a depth of 1380 ft. Two exploration levels have been driven within the ore zone at depths of 165 and 1246 ft below the surface outcrop. The development to date had the objective of sampling the mineralization and produc¬ing detailed information on the outline of the ore body and the distribution and controls of zinc and lead values. In an attempt to satisfy the basic design criteria for the mine, it was decided that production would be best achieved by a combination of 40% sublevel long-hole stoping and 60% cut-and-fill mining. Costs of exploration and capital development of per¬manent underground facilities are normally written off over the life of a mine. Production expenditures, on the other hand, are of a temporary nature and are normally charged as and when incurred as an operating expense. Reasonably accurate predictions of mine production costs can be built up from engineering design and estimates of individual mine activities for ultimate inclusion in the comprehen¬sive data required for financial decision making. The simulated operations can be costed on a detailed basis in the form of a monthly operating budget. The budget format can be generalized or detailed, depending upon the scope of the project. However, ex¬perience suggests that a fairly detailed format has the advantage of assuring that all significant cost items are included. For underground costing it is suggested that the budget structure include five major cost centers (i.e. development, diamond drilling, ore extraction, hoisting/ transportation, and general mine expense). These in turn are detailed under numerous subheadings. The mechanism for compiling an operating budget will be illustrated. Because of its relative simplicity, ore extraction under sublevel long-hole stoping has been chosen for illustration. All other activities, simple or complex, can be estimated in similar fashion. BLOCK AND STOPE DEVELOPMENT Long-hole blocks, used where advantageous, will be up to 250 ft in height, depending upon the vertical con¬tinuity of the mineralization, and approximately 300 ft long. Drawpoints will be at 36-ft intervals and serviced by loading crosscuts driven from a footwall drift parallel to and close to the ore zone. Pillars between the stopes will be 50 ft wide. Stopes will be drilled off with vertical rings of blastholes drilled from sublevels approximately 60 ft apart vertically. This drilling will be done by percussion drilling machines (31/2 in.) mounted on a trackless drilling rig. Load¬haul-dump (LHD) equipment will be used to move broken ore from the drawpoints to the orepass connecting to rail haulage systems. On completion, long-hole stopes will be backfilled to prevent caving and to facili¬tate later pillar removal. From a planned stope layout, a forecast of produc¬tion and development is made in Table 1. Table 1. Block Tonnage and Stope Development Quantity Ore Waste Total ore block 375,000 st 2 stopes 310,000 st 1 pillar 65,000 st Access crosscuts, 4 at 100 ft 400 ft Drill sublevel drifts, 6 at 300 ft 1800 ft Stope raises, 3 at 250 ft 750 ft Undercut sublevel drifts, 2 at 300 ft 600 ft Loadout crosscuts at 35-ft intervals 550 ft 100 ft 3300 ft 500 ft Total development footage 3800 ft Tons per ft of development 987 st

Jan 1, 1982

By R. Woods, J. D. Wright

INTRODUCTION The Endako deposit was first explored in 1926, but development waited until the early nineteen sixties. It first came under the control of Placer Development in 1963, and early in 1964 a decision to bring it into production was reached. The deposit is located near the geographic center of British Columbia, and is served by an east-west communication system including a paved highway, mainline railway and good telephone, teletype, and computer facilities. The plant area is shown in Figure 1. The operation is known as the Endako Mines Division of Canex Placer Limited of Vancouver. First production was achieved in June, 1965, at a nominal rate of 10,000 tons per day, but it soon was closer to 15,000 tons per day. In 1967, two new grinding and flotation sections brought capacity to the present rate of approximately 27,000 tons per day. With mill heads of around 0.15 percent MoS2, production is about 14,000,000 pounds of molybdenum per year. A flowsheet of the operation is given in Figure 2. A legend for this flowsheet is given in Table 1. Published ore reserves are estimated at 214,500,000 tons at an average grade of 0.143 percent MoS2. PRIMARY CRUSHING The 85 or 100-ton haulage trucks dump over a grizzly into a 42 inch x 65 inch C.A.C. (Canadian Allis Chalmers) gyratory crusher.

Jan 1, 1976

F. E. Bachman, Port Henry, N. T. (communication to the Secretary*):—In discussing Mr. o.o.Laudig's paper, the Action of Blast-Furnace Gases Upon Various Iron-Ores,' I took the ground that Mesabi-ore explosions were entirely mechanical; but more observation and later experience has caused me to

Jan 1, 1905

By Frank E. Chase

Retreat mining is one of the most hazardous operations conducted underground. One of the keys to miner safety and an efficient recovery of the reserves is to design adequately sized production pillars to prevent pillar squeezes, excessive pillar spalling, severe floor heave, roof falls, and pillar bumps. Currently, few pillar design approaches are available which consider the abutment pressures generated by the gob. The Bureau of Mines is developing a method called Analysis of Retreat Mining Pillar Stability (ARMPS) to aid in the design of retreating sections. Back analysis of 42 case studies indicates that ARMPS can be helpful in predicting stability during retreat mining operations.

Jan 1, 1993

The silver-lead ores of Yerranderie occur in a series of narrow, flat-dipping lodes in small bodies of quartz porphyry, intrusive into Upper Devonian volcanics, which are exposed in the western flank of the Permo-Triassic Basin of central New South Wales, about 80 miles south-west of Sydney (Harper, 1930; Lawrence, 1953). The lodes strike about E.-W. and are from 1,200ft. to 2,000ft. long. They consist of a series of ore shoots from 3ft. to 100ft. long, and up to 5ft. thick, though commonly under 2ft. thick, interspaced with low grade sections. They dip at from 18°S. to 400 S., and have persisted down the dip for 1,450ft. (693ft. vertical depth) in the Wollondilly mine, for 1,120ft. in the Bartlett's mine, and for 850ft. in the Wonga mine.The grade of ore in the Bartlett's mine was from 32% to 40% lead, from 100 oz. to 130 oz. silver per ton, 0.5 to 4.0% arsenic, with ore ,as rich as 20% arsenic on the No. 6 and No. 7 levels of the Silver Peak mine, the underlay extension of the Bartlett's lode, about 1% copper, and up to 4 dwt. gold per ton. In the Wollondilly mine the average grade was 15% lead, with sections as rich" as 36% lead, from 75 to 165 oz. silver per ton, 0.6 to 3.3% arsenic, 0.2 to 0.7% antimony, 0.8% copper, and up to 7...5. Cl~t. gold per ton. The zinc content of the ores is slight, and has not been recorded.Mining began about 1898, and to date more than 110,000 tons of ore, valued at £1,800,000, has been raised. Large scale mining ceased about 1935. The chief mines operating were the Bartlett's mine (formerly Colon Peak), the Silver Peaks, which is the underlay extension of the Bartlett's, the Wollondilly (or Yerranderie) mine, and the Wonga' (or Tonalli) mine.

Jan 1, 1953

By M. S. Seltzer

High temperature steady-state creep rates have been determined jor lead sulfide single crystals whose defect concentrations were fixed by equilibration under controlled sulfur pressure. The activation energy of creep, Qc , for undoped lead sulfide, is a function 01 composilion, decreasing from 2.9 to 2.2 ev as the carrier concentration increases from 9 x 1016 to 8 1018 electrons cm-3 for lead excess crystals, and increasing from 1.7 to 2.3 ev as the concentration of holes increases from 4 x 1017 to 1 x 1018 cm-3 in sulfur excess material. For five different composilions, Qc for undoped and bismuth-doped crystals has been found to be equal to Qpb + Qs, the self-diffusion activation energies for lead and sulfur in PbS. It is well-established1 that high-temperature steady-state creep of pure metals is a diffusion-controlled process, where the activation energy of creep is equal, within experimental error, to the activation energy of self-diffusion. The relationship between the activation energy for steady-state creep and those for self-diffusion in binary compounds is more complicated, however, because two different diffusing species must be considered, and becayse the activation energies may be influenced by compositional variations and by foreign atom additions.2-4 In some cases experimentally determined creep activation energies are correlated with those for self-diffusion of the slower-moving species, but such agreement is by no means universal. Interpretation of experimental results is difficult because few investigations include measurements of creep rates and self-diffusion coefficients of both diffusing species on single crystals of similar composition and foreign atom concentration. The objective of this investigation, therefore, is to determine the influence of stoichiometric defects and foreign atoms on steady-state creep and self-diffusion for a well-characterized binary compound. Lead sulfide, a compound semiconductor which crystallizes in the NaC1-type structure, and whose homogeneity range includes compositions on either side of the 1:l atom ratio, was chosen for study in this investigation. The pressure-temperature-composition relations for PbS are available,5,3 and deviations from stoichiometry may be monitored by electrical measurements.7 Single-crystal specimens are available to allow observation of effects due to bulk phenomena without interference from grain-boundary effects. These crystals have been employed to determine steady-state creep rates under compressive stress for "pure" and bismuth-doped PbS as a function of sulfur pressure and temperature as well as crystal orientation and applied stress in order to establish the relationships between defect concentrations and creep rates. EXPERIMENTAL PROCEDURES Crystal Growth. Lead sulfide crystals, prepared by the Bridgman method, are grown in evacuated quartz crucibles, filled with almost equal atomic proportions of high-purity (99.999 pct) lead and sulfur, and heated to a temperature above the melting point of PbS (about 1127°C) so that the lead and sulfur react to form molten PbS. The crucible and melt are dropped through the hot zone of the crystal-growing furnace at the rate of 1 cm per hr. The maximum melting point of PbS is slightly on the lead-rich side of the 50:50 composition; hence crystals grown from a near stoichiometric melt are usually n-type semiconductors containing excess lead atoms. Typically the lead excess concentration in solid solution with the PbS is 3 x 1018 atoms cm-3. Crystals containing either 1/20 or 1/2 mole pct Bi2S3 were prepared by addition of the foreign atoms prior to crystal synthesis. Preparation of Specimens of Controlled Stoichiom-etry. Since it is difficult to grow crystals of desired composition, the process of crystal growth and control of stoichiometry are separated into two distinct operations. Equilibration of as-grown PbS to a desired composition is most easily accomplished by first annealing the n-type crystals under high sulfur pressure (about 1 atm) at temperatures above 700°C. This process serves to put excess lead precipitates into solution, yielding 0 -type crystals upon quenching to room temperature. Inter diffusion occurs readily in crystals so treated, and the methods described by Bloem may then be employed to obtain crystals of controlled stoi-chiometry. For the creep measurements made in this study, this involves heating crystals in mixtures of H2 and H2S The equilibrium constant for H2S is given by: Thus the sulfur partial pressure is obtained from knowledge of the H2/H2S ratio and Kh2S for the temperature in question. Fig. 1 shows the p-T-x diagram for PbS. It can be seen that, within the stability limits for single-phase lead sulfide, desired compositions may be obtained by equilibrating under a specified sulfur pressure and temperature. Characterization of Crystals. The principal methods for evaluating PbS crystals include measurement of electrical properties and mass-spectrographic analysis. The measured electrical parameters include Hall coefficient, conductivity type, and electrical resistivity Bloem7 has shown that there is a 1:l relation between the defect concentration and carrier concentration obtained. Hence, if the contribution to the carrier con-

Jan 1, 1968

By John A. Herbst

Professor A.M. Gaudin and his students at MIT were dedicated to understanding the fundamentals of mineral processing and as a consequence were constantly searching for better quantatative descriptions of size reduction, liberation and mineral separations. I have had the good fortune of meeting Professor Gaudin and working with many the modeling pioneers who followed him. I have observed first hand the refinement of mineral processing models as they increased in realisticness moving up the steps of the pyramid (see figure below) from pure empiricism to the current pinacle of physics based models. During this journey our industry benefited significanbtly as model prediction errors were reduced from ± 20% to ± 1%. Unit operation and plant simulations improved in accuracy and speed of calculations.

Sep 1, 2012

By F. Bourgeois, J. -S. Kroll-Rabotin

ABSTRACT: Falcon concentrators are enhanced gravity separators designed for concentrating fine particles. The Falcon UF model is unique in that it is dedicated to beneficiation of ultrafines, one key feature being that it does not make use of any fluidization water. We investigated the physics of particle transport inside Falcon concentrators, and concluded that separation efficiency is governed by differential settling velocity. We derived and published a predictive model of the partition function under dilute conditions. We intend to extend the initial model to concentrated ultrafine suspensions for application to industrial scenarios by adding hindered settling to account for solid concentration effects.

Jan 1, 2011

By A T. Spathis, K E. Leslie, K R. Blay

Many open pit mining operations seek to control blasting-related movement of ore and waste to minimise loss and dilution. Tracking the movement of ore boundaries can result in the maximisation of ore recovery and minimisation of waste sent to the mill, thereby optimising the economic performance of the operation.A novel solution for tracking the ore–waste boundary location before and after a blast uses subsurface markers placed along the boundary with all measurements and scans conducted remotely from the blast block. Upon system activation, user-operated equipment on the surface determines the three-dimensional (3D) location of the subsurface markers.The patented system concept combines two electromagnetic mechanisms, magnetic positioning and radio frequency through-rock communications to enable remote battery-powered markers, buried within the rock to be accurately located in 3D space.Field trials at a quarry have demonstrated the system can comfortably handle over 100 subsurface markers with a complete scan, including set-up, taking less than an hour. Location accuracy for each marker in 3D space is better than 0.5 m at a range up to 50 m from the magnetic source, and the system has an overall operational range in excess of 75 m.CITATION:Blay, K R, Spathis, A T and Leslie, K E, 2015. A remote, rapid and accurate three-dimensional ore waste boundary tracking system, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 297–306 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By Valerio Violo, Antonio Raschillà, Livia Cicinelli

"The Cityringen project foresees the construction of a new circular metro line in and around the city center of Copenhagen. Both drives, the north and the south one, required the TBMs to be equipped in order to cope with contaminated water table and soil. The north drive mined in water contaminated with chlorinated compounds, including chloroform and vinyl chloride, while the south TBMs excavated in a highly benzene contaminated area, becoming the first TBMs to complete such a challenge. Despite the pollution and the stricter operative procedures, the production has been extremely successful with record up to 520 m in one month and 32 m in one day. 1 INTRODUCTION The Cityringen is an extension of the existing metro line. The 15 km long twin tunnels will form a new circular line with 21 new stations (17 + 4 crossovers). When finished it will interchange with the existing metro, rail and bus infrastructure at key locations around town, taking 90% of Copenhagen citizen at less the 600m from a metro station (Figure 1). The Contract has been awarded in April 2011 to the Copenhagen Metro Team (CMT), a joint venture between the Italian companies Salini Impregilo, Tecnimont and Seli SpA. Tunnels are being excavated with 4 Seli-Kawasaki EPB TBMs, two of which manufactured in Denmark by Seli Tunneling Denmark ApS, which is also operating them. At the moment this article is written 50% of the tunnels have been completed and over 40% handed over for the electromechanical works. Seli Tunnelling Denmark is a Danish company constituted in 2012 as a subsidiary of Seli SpA for this project and subsequently acquired by the Salini Impregilo Group in 2015."

Jan 1, 2016

By Robert A. Irvine

Traditionally, cost effective gains in the capacity of material handling systems were a result of increases in equipment size. In general, the use of maximum capacity components in a system optimizes the production "costs per ton." The trend-"Bigger is Better" will continue in material handling equipment. Since it is rare for plant operators to lament an over abundance of screening capacity, one can safely predict that vibrating screens will follow this trend. Large Vibrating Screens - The Challenge Large vibrating screens represent a unique challenge for Manufacturers, Plant Designers, and Plant Operators. The inherent mode of operation for vibrating screens is self-destructive. More often than Manufacturers admit, Designers plan for, or Operators staff for, a vibrating screen succeeds and self-destructs. This is a problem. It can magnify with larger vibrating screens.

Jan 1, 1983

By A Harburg, A Dendys, G Green

Dugald River Mine (DRM) is located approximately 85 km north-east of Mount Isa, and 65 km northwest of Cloncurry, in the north-west Queensland Minerals Province of Australia. DRM is an underground operation and includes an above ground processing area, a 200 ha tailings facility, two waste rock dumps and associated infrastructure. The mining method is conventional underground longhole open stoping and downhole benching of a lead/zinc deposit within a black slate environment. The deposit is estimated to contain ~53 Mt of zinc and lead, which will be mined over a 30-year period. During operations, waste rock is segregated into either Potentially Acid Forming (PAF) or Non-Acid Forming (NAF) based primarily on the sulfide concentration of the rock. In particular, the NAF waste rock is an asset and is used for various small construction projects during operation; however, its primary use is for rehabilitation work at closure. Since beginning operations in 2017, DRM has generated more NAF than was originally anticipated. Whilst this is an excellent asset to DRM, it did pose a logistical problem of where to store the additional NAF for the life-of-mine and how to manage the run-off. DRM experiences distinct wet and dry seasons with hot conditions and periods of rainfall between November to April, and relatively dry and mild conditions between May to October. Majority of rainfall is received between January and February with an annual average rainfall of 501.1 mm. Short and intense rainfall events are common. DRM’s above ground infrastructure is land-locked between two large ephemeral watercourses, making the significant expansion of any assets challenging from a permitting/regulation and cost perspective. MMG is also a member of the International Council on Mining and Metals (ICMM) and is committed to adapting to a changing climate through operations and closure. Following extensive studies DRM opted to store the additional NAF material into the adjoining NAF Dam and concurrently upgrade a previously minor sediment basin into a highly engineered sediment basin which relies on automatic flocculant dosing to meet the required discharge quality limits. This basin is climate-resilient in that the dosing system can be adjusted to counter high intensity extended rainfall events which produce highly variable inflow water quality. This abstract broadly describes how DRM designed and implemented an engineering innovation to enable the site to meet the changing climatic challenges. The abstract then describes the process to design the sediment basin and how the dosing rate was established and validated. Finally, the abstract comments on the effectiveness of the sediment basin and how it has revolutionised DRM’s approach to water treatment and management; all whilst minimising the disturbance footprint of the operation

Aug 2, 2023

By P. G. Rueffel

"Natural deposits of sodium sulphate are the result of normal geological processes. They are world-wide in occurrence and are found in depressions in arid and semi-arid regions. Primarily, the sodium sulphate deposits of the prairies are post-glacial and differ from rock salt deposits, which originated from connate water (fossil sea-water). Sodium sulphate on the Canadian prairies occurs as brine or in beds of intermittent and/or permanent crystals. Most beds are contaminated with various amounts of silt, sand and decayed organic matter. Soluble impurities such as chlorides, carbonates and bicarbonates of sodium and sulphates of magnesium and calcium are leached out from the soil of the surrounding run-off area of the deposits.In this paper, the origin of sodium sulphate, and the physical and chemical aspects related to the development of the Ingebrigt Lake deposit are dealt with. The problems of garnering raw material, and of refining and de-hydration are discussed, with emphasis on the Ingebrigt Lake plant. The utilization of the ""Total Power Concept"" in a sodium sulphate plant is described. The principal uses of sodium sulphate are also discussed."

Jan 1, 1968

The operations management system at Olympic Dam has evolved over a two-year period to the state described in this paper. This evolution has been driven largely by the requirements of the recent expansion of the Olympic Dam operation from three to nine million tonnes of ore per year. Such a change process demands an holistic approach to mine management and several techniques have been employed including XPAC, Mine Works Planner as well as a basic philosophy of æjust-in-timeÆ completion. The philosophies, software and tracking mechanisms for managing a large tonnage operation are described, outlining a fully integrated scheduling system from shift to shift detailed activity planning to a 20 year production plan. Moreover, the vision of production exceeding the current limit of nine million tonnes means that there are to be further developments to this state-of-the-art system in use at the Olympic Dam mine.

Jan 1, 2000

By R. D. Currie

Many coal companies have a system for checking men into and out of the mine , and generally such systems appear to be perfectly sound and adequate . However , when the real test comes at the time of a mine disaster , too often the system is found to have failed , and hours and sometimes days pass before it is known how many men are in the mine and who they are . Most of the systems in use , when properly maintained , are adequate for the routine work of the mine , and some of them may prove adequate at the time of a mine disaster . All systems , however , depend upon proper supervision and strict adherence to the system .

Jun 1, 1930

By Dao Su, Ganghua Fu, Xuling Chen, Yufeng Guo, Min Gan, Tao Jiang, Xiaohui Fan

"The sinter quantity and quality of limonite in the magnetite and hematite sintering models were studied and compared. For the sintering of magnetite, the results showed that the increase of limonite improved the permeability of sinter mixture, increased the sinter quantity and quality. The sinter strength and productivity were decreased when the proportion of limonite exceeded 50%. The appropriate proportion of limonite was 40%-50%. For the sintering of hematite, the sinter quantity and quality were first maintained at the same level thereafter declined, more obviously when the proportion of limonite exceeded 40%. The amount of calcium ferrite increased whose shape changed from acicular to leaflike and block gradually with the proportion of limonite increased from 0% to 80%.IntroductionWith the fierce competition of domestic and overseas iron markets in the twenty-first century, cheap ores such as limonite are used to be sintering raw materials to cut the costs of sinter and pig iron for steel making, and boosts the competitiveness of steel companies [1, 2]. In addition, as the reserve volume of hematite of high grade is reducing year by year, the usage amount of limonite in sintering are expanding in major domestic and international sintering plant [3, 4]. Sintering can be divided into two types, majoring in concentrates and fine ores, according to the different types of raw materials. The sintering characteristics of limonite in the two types above have been studied and compared in this paper, which makes a significant sense for the research of limonite sintering properties and guidance of ore proportioning in sintering plants."

Jan 1, 2012

By P. J. Shenon

ENROLLMENT in geological and mining engineering curricula is declining at an accelerated rate despite the greatest need for trained men ever extant in the minerals industry. Industrial and military demand is mounting, but the number of freshmen selecting the mineral field continues to fall. Estimates on the needs of industry range as high as 30,000 new engineers a year. The current deficit is more than 60,000 engineers less than the 350,000 to 450,000 which eventually will be needed. The indisputable fact is that the colleges are turning out fewer and fewer engineers despite the greatest enrollment in colleges and universities ever experienced in the United States. In 1950 a record 52,000 young men stepped out of the confines of ivy covered walls with engineering degrees in their hands. By 1951, however, the number dropped to 41,000 and present enrollment indicates a national graduating class of only 25,000 for 1952. No letup in the drop is forecast. About 19,000 can be looked for in 1953 and 1954 may reach an unhappy 12,000. It becomes clear that something must be done to attract high school graduates to engineering. One immediate possibility could be to make the course burden carried by the engineering student somewhat lighter. The prescribed curriculum in many schools is such that the student takes the path of least resistance, and instead of training for an engineering future, studies for a vocation which will allow him to learn and at the same time get at least a nominal enjoyment out of college life. Review geological and mining curricula of 20 colleges and it will be found that the engineering student is a veritable pack mule compared to a lad taking liberal arts or some other non-technical program of study. The curriculum for geological engineering at one school calls for 202 semester hr, with almost 23 hr carried per semester. Multiply this figure by three hr, the minimum supposedly to be devoted to a credit and you get 69 hr per week. With a bare minimum of 84 hr for sleeping and eating, about two hours a day remain for recreation. However, the load of other schools investigated is about 19 hr. The University of Utah requires 238 quarter hr for graduation with a degree in geological engineering, while requiring only 183 quarter hr for baccalaureate degree from University college, Utah's liberal arts school. It can be stated with a measure of surety that the same proportions exist in other universities. The first step would be for ECPD to review its requirements for mining and geological engineering. It must recognize that mining and geological engineers operate in a specialized field, as do other types of engineers. Although a geological engineer may not design a bridge, as pictured by the ECPD Committee on Engineering Schools, his field of design calls for similar engineering precision, a knowledge of materials, construction methods, economic considerations, and financing. Six schools have been accredited by the ECPD. What is the basis for approval and can the requirements be modified and still be kept in line with the needs of the geological engineer? Course work from school to school varies with the exception of mathematics, chemistry, and physics. Even in those courses the not inconsiderable variation lends dubious creditability to the mean. One accredited school requires 7 1/3 semester hr of chemistry, compared with 24 hr required by another, making an average for the six schools of 17 1 /3 hr. Required credit hr in mechanics ranges from 4 to 18 and in surveying from 2 to 15. Several non-accredited schools require more hr than do the accredited schools in some courses. Why is the engineering student forced to carry such a back-breaking load? The answer is of course fairly obvious. He is irrevocably set apart from the rest of the student body because of the nature of his life's work. He is training for a place in a world where technology is becoming increasingly involved. He must be prepared to do a job now-and not later. Mining and geological engineering require the same essential backgrounds as other engineers, and more. The "more" is a knowledge of mining methods, metallurgy and geology for the mining engineer. The geological engineer must know in addition, mineralogy, petrography, and geophysics. The load is compounded finally by the addition of liberal arts courses. Should anything be done to relieve the situation? Today's engineer must be a whole man, capable of handling the tools of communication and with an understanding of the economics of industry. He must be able to write clear simple English, and he must be man who can think from some other position than bent over a work table. He must be aware of the history of his country and to some extent that of the world. Not all schools share this view. Only two of the accredited schools require history courses. However, five of the non-accredited schools make it mandatory. Four accredited and five of the nonaccredited schools require economics. Courses in mathematics, physics, and chemistry are fundamental in engineer training. The average for the accredited schools could serve as a guide in

Jan 1, 1952

Council and Standing Committees, 1976-1977 President: Dr M. D. G. Salamon. Vice-Presidents: P. A. von Wielligh and Dr M. G. Atmore. Honorary Treasurer: J. K. E. Douglas. Immediate Past President: Dr R. E. Robinson. Members of Council: J. D. Austin, E. Collier, C. T. Fenton, J. A. Holmes, Dr P. R. Jochens, Prof. R. P. King, Dr D. G. Krige, Prof. F. Q. P. Leiding, D. G. Malan, C. T. Shaw, D. A. Viljoen, R. A. von Maltitz. Chairman, Witbank-Middelburg Branch: E. J. Jackson. Chairman, Orange Free StatelKlerksdorp Branch: A. N. Shand. B. Past Presidents Serving on Council: R. J. Adamson, H. Britten, J. K. E. Douglas, R. C. J. Goode, Prof. D. D. Howat, Dr J. P. Hugo, D. G. Maxwell, Prof. R. P. Plewman, V. C. Robinson, P. W. J. van Rensburg. Ten Council meetings were held during the year, and the standing and sub-committees held seventy-one meetings. Standing Committee Chairmen: Membership, Dr M. G. Atmore; Technical Programme - Metallurgy, J. A. Holmes ; Technical Programme - Mining, C. T. Shaw; Vacation School - Metallurgy, Prof. R. P. King ; Vacation School - Mining, Dr D. G. Krige ; Awards, Dr R. E. Robinson; Excursions, C. T. Fenton; Adjudi¬cation, Dr R. E. Robinson; Constitution, Rules and Policy, Dr R. E. Robinson; Mines and Works, C. T. Shaw; Education, Dr R. E. Robinson; Pyrometallurgical Discussion Group (Convenor), Dr J. B. See; Phoenix National Committee, Dr R. E. Robinson.

Jan 1, 1978

Bailey, Lewis Newton, Master Engineer, Senior Grade, 4th Regiment, U. S. Engineers, Headquarters Company, died of pneumonia at Camp Merritt, N. J., on April 30, 1918. Baird, Louis, Lieut., Royal Field Artillery, British Army, died on the battlefield in 1915. Cobeldick, William Morley, Royal Engineers, died from gas poisoning on October 7, 1915. Dougall, Ralph, killed in action. Evans, Alfred Winter, Lieut.-Col., New Zealand Rifle Brigade, D. S. 0., D. C. M., killed in action on October 12, 1917. Gorman, Thomas C., Lieut., Canadian Engineers, killed in France, Mar. 18, 1918. Hague, William, 1st Lieut., Engineer Officers' Reserve Corps, died in active service. Hall, William T., Capt., Royal Flying Corps, killed in action. Heine, Bernhardt E., Lieut., Aviation Service, killed at Fort Sill, Okla., Aug. 10, 1918. Irving, John Duer, Capt., 11th Engineers, A. E. F., died while on active service in France on July 26, 1918.

Jan 9, 1918

By Tapp BA

Australia, considered an energy and mineral resource rich country, is poised on the verge of a major mineral industry 'boom'. It has been repeatedly emphasised that this growth in the mineral industry will form the basis of future economic development. Currently the mining and mineral processing industries use some 32% of the annual primary energy consumption in Australia. This share, although large even now, will increase with the realisation of the 1980's production expansions to 40% by 1985. Two metal industries dominate the energy scenario; iron and steel and aluminium. Major expansions in the local industry will be dominated by the aluminium industry. This highly energy intensive industry has planned major increases in metal production throughout the eighties. The implications of exporting cheap energy to pay for expensive oil imports, the ramifications on Australia's infrastructural ability to supply these major energy needs and the other options that might be

Jan 1, 1981