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Two examples of risk assessment and risk management in action are discussed. The first is in a management context in relation to Kingsgate Consolidated LtdÆs Chatree operation in Thailand. Kingsgate has conducted an over-riding risk management program to assess and manage risk across its complete business from the Sydney and Bangkok offices through to the Chatree Mine and the companyÆs exploration group. This example addresses how Kingsgate has applied risk management, particularly to the process operation, and the benefits that have been gained. The second is in an operational context in relation to Minara Resources LtdÆs Murrin Murrin operation in Western Australia. The subject is the methodology and outcomes of a risk assessment across the Murrin Murrin operation specifically to identify and assess risks with potential impacts from $1 M to $7 M. This section of the paper then shows how the outcomes of the risk assessment were integrated into MinaraÆs strategic plan to debottleneck the plant and increase nickel production.

Jan 1, 2004

Council Members 1988/1989, left to right: Row 1 (seated): H.E. James, P.W.J. van Rensburg, Dr H. Wagner, B.C. Alberts, C.E. Flvaz (President), H.G. Mosenthal, R.D. Beck, Professor A.N. Brown, Dr P.R. Jochens Row 2 (standing): P.C. van Aswegen, J.P. Hoftman, D. Wllson, P.M.T. White, K.A. van Gessel, Dr LA. Cramer, J.S. Freer, V.J. Moore, G.A. Brown, R.A. Snodgrass, J.D. Austin Row 3 (standing): Dr G.A. Fourle, Dr J. Lurle, A.M.P. Henderson, R.P. Mohring, J.A. Cruise, J.L. Nel, H. Scott Russell, J.P. Deetlefs, D.A.J. Ross-Watt Inserts, top row: K.R. Beard, N.J. De~ine, R.J. Dlppenaar, G.T.G. Emere, Dr J-P. Franzidls, Professor R.P. King, J.G. McCallum. Bottom row: G.Y. Nisblt, Professor R.P. Plewman, Dr R.E. Roblnson, Dr O.K.H. Steften, P.A. van Wlelllgh

Jan 1, 1989

By William Ewing

Copper is rapidly being adopted by the semiconductor industry as the interconnect material of choice. The aqueous processing techniques used generate wastes such as spent electrolyte from electroplating, electroplating rinse water, copper solutions from removing copper from the back of the wafer and CMP waste streams. We are examining the use of chelating resins as a means of recovering copper from these streams within the processing plant, thereby allowing recycling of process water and minimizing hazardous waste disposal costs. We have examined the effect of model plating additives (and other organic complexing agents) on the efficacy of Dowex M4195. Although these might have had a significant effect, commonly used electroplating additives polyethylene glycol (PEG), benzotriazole (BTA), thiourea, glycine, and 3-mercapto-1-propanesulfonate (MPSA) each had little or no effect on the ability of the resin to chelate copper and remove it from solution. This suggests that Dowex M4195 could be useful for treating waste solutions of rather variable compositions.

Jan 1, 2003

By E. Topuz, M. J. McPherson, I. M. Konduri

Theoretical investigations and laboratory experiments demonstrated that there is a great potential for the use of jet fan ventilation in mines. This paper summarizes research that consisted of laboratory-scale model tests, full-scale mine model tests and underground coal mine entry tests. The laboratory model used water as a fluid medium and simulated a typical 25-m- (82-ft-) long heading having a 1.5 x 6-m (4.9 x 19.7 ft) cross section. The effects of various nozzle positions and designs were investigated by using still photography. It was also found that the use of jets in combination with line curtains enhanced the overall ventilating effect in the model entry. Flow measurements in the full-scale mine model and in the coal mine entry were conducted to evaluate the effectiveness of the jet fan. It was concluded that the designed jet fan ventilation system was an effective means of ventilating blind airways.

Jan 1, 1998

Book Review Gold Metallurgy in South Africa, edited by R. J. Adamson and published by the Chamber of Mines of South Africa, Johannesburg, 1972. The metallurgical processes used for the recovery of gold from Witwatersrand ores were first described in A Text Book of Rand Metallurgical Practice, published in 1912 and subsequently revised in 1926. This was followed by Gold M etall1trgy on the Witwatersrand by A. King in 1949. New Publications Elements of Practical Coalmining, S. M. Cassidy (editor), New York, Society of Mining Engineers of AIME, 1973, $16.50. Future Enerey Outlook, J. A. Pederson (editor), Colorado School of Mines Quarterly, vol. 68, no. 2, April 1973, $6.00. Guide To The Preparation Of Papers For Publication In The Journal Of The South African Institute Of Mining And Metallurgy The following notes have been compiled to assist authors in the preparation of papers for presentation to the Institute and for publication in the .Journal.

By A E. C Peres, J D. Donda

With the aim of producing a concentrate with silica content lower than 1.0 per cent SiO2, laboratory experiments were conducted using a sample of the final concentrate from the Samarco iron ore flotation plant (mechanical cells circuit). The sample contained 3.08 per cent SiO2. This sample was split in two size ranges: >105 ¦m and 2. The >105 ¦m fraction, representing 23.3 per cent of the initial mass, was re-ground to 100 per cent 2 content were obtained. Considering the mass proportions of the two size fractions, the calculated silica content of the combined concentrate was 0.68 per cent, which is considered satisfactory for the objective of this investigation. The major reasons for the higher (0.90 per cent) silica content of the concentrate from the coarse fraction are the presence of silicon in the lattice of goethite and the presence of residual silicious slimes in the pores of porous hematite.

Jan 1, 2005

(Members are urged to send in for this column any notes of interest concerning themselves or their fellow-members.) Members and guests who registered at Institute headquarters during the period April 1 to 15. William B. McKinley, New York, N. Y Collier Cobb, Chapel Hill, N. C. C. W. Goodale, Butte, Mont. E. P. Fleming, Guayacan, Chile. D. C. Brown, San Luis Potosi, Mex. Albert Sauveur, Cambridge, Mass. W. H. Rettie, Yonkers, N. Y. H. M. La Follette, La Follette, Tenn. Mark B. Kerr, San Francisco, Cal. E. Fleming L'Engle, New York, N. Y. H. W. Turner, San Francisco, Cal. H. Garlichs, St. Louis, Mo. Frank M. Leland, Mackay, Idaho. Henry Hay, London, England. Sir Robert Hadfield, London, Eng. Edward J. Way announces that his agreement as consulting engineer in South Africa for the Anglo-French Exploration Co., Ltd., having terminated, he will engage in professional consulting work with headquarters at 135 Cullinan Buildings, Simmonds St., Johannesburg, South Africa.

Jan 5, 1914

By K. Koyama

As a study to establish a new energy-saving copper recycling process, the copper deposition at a cathode and oxidation of copper(I) at an anode from ammoniacal alkaline solutions containing Cu(I) or Cu(II) were investigated in order to examine the feasibility of electrolysis in this solution.. The current efficiencies were greater than 95% in the current density range of 200 to 1000 A m-2. The current efficiency decreased with the increasing Cu(II) concentration and temperature. In the batch-type electrolysis at the anode, during the initial stage, the current efficiencies were 100%, and the anode potentials ranged from 0 to 0.15V vs. SHE at a current density of 100 A m-2. During the subsequent stage, the current efficiency decreased with time, and the anode potential was around 1.35V vs. SHE where oxygen evolution at the anode surface was observed. As a result, the present electrolytic method can be the core technology in the said copper recycling process.

Jan 1, 2007

COUNCIL AND STANDING COMMITTEES, 1974-1975 Office Bearers : Prof. R. P. Plewman (President), Dr R. E. Robinson and Dr M. D. G. Salamon (Vice Presidents), J. K. E. Douglas (Honorary Treasurer), and P. W. J. van Rensburg (Immediate Past President). Elected Members of Council: Dr M. G. Atmore, E. Collier, G. H. Grange, Dr P. R. Jochens, Dr D. I. Legge, D. G. Malan, C. T. Shaw, L. W. P. van den Bosch, D. A. Viljoen, H. M. Wells, P. A. von Wielligh. Branch Chairmen: B. T. Hosking (Witbank-Middelburg Branch), D. A. Smith (Orange Free State Branch). Past Presidents Serving on Council: R. J. Adamson, M. Barcza (until May 1975), H. Britten, H. E. Cross, Prof. J. de V. Lambrechts (until December 1974), R. C. J. Goode, Prof. D. D. Howat, Dr J. P. Hugo, D. M. Jamieson, D. G. Maxwell, Dr J. T. McIntyre, V. C. Robinson. Ten Council meetings were held during the year, with an average attendance of 19, and the standing and sub-committees held 74 meetings. COMMITTEE CHAIRMEN Membership, Dr M. D. G. Salamon; Technical Programme-Metallurgy, D. A. Viljoen; Technical Programme-Mining, G. H. Grange ; Vacation School-Metallurgy, Dr P. R. Jochens; Vacation School-Mining, Dr M. D. G. Salamon; Awards, Prof. R. P. Plewman; Excursions, Dr D. I. Legge; Constitution, Rules and Policy, P. W. J. van Rensburg and Prof. R. P. Plewman; Mines and Works, P. A. von Wielligh; Education, P. W. J. van Rensburg. REPRESENTATIVES ON OTHER BODIES Associated Scientific and Technical Societies Members: V. C. Robinson and Prof. R. P. Plewman. Alternates : Dr R. E. Robinson and Dr M. D. G. Salamon. Federation of Societies of Professional Engineers Members : J. K. E. Douglas and P. W. J. van Rensburg. Alternates : Prof. R. P. Plewman and G. H. Grange. South African Council for Professional Engineers Member : R. C. J. Goode. Alternate : G. H. Grange. Engineers' Liaison Committee (Pretoria) Member: Dr J. P. Hugo. Alternate: P. W. J. van Rensburg. MEMBERSHIP Thirteen Fellows, seventeen Members, eleven Graduates, three Associate Members, eighteen Associ¬ates, twenty-six Students, and one Company Affiliate were elected during the year. Nine Fellows, four Members, and three Associates were admitted to retired membership. Eight Fellows were elected to Honorary Life Membership. Three Members and one Graduate were transferred to Fellows, nine Graduates and two Students to Members, six Students to Graduates, one Student to Associate, and one Associate to Associate Member. Six Fellows, ten Members, three Associates, and one Student resigned. A statement of membership as at June 30th 1975 and the end of the previous year is shown in the following tabulation: 30th - 30th June, - June, 1974 - 1975 Honorary Life Fellows - 9 - 15 Honorary Fellows 8 - 8 Life Fellows 170 - 170 Fellows 544 - 536 Members 512 - 523 Associate Members 10 -15 Graduates 130 - 133 Associates 194 - 210 Students 173 -175 Company Affiliates 70 - 71 1 820 - 1 851 [ ] OBITUARIES Your Council records with deep regret the death during the year of the following members: M. Barcza, Honorary Life Fellow and Past President of the Institute; J. A. Boyd, F. E. Keep and J. H. Taylor, Life Fellows; I. D. B. Corner, C. O'G. Deane, R. E. Gilmour, A. C. Hofmeyr, M. W. Howell, T. J. Robin and F. C. Steinhobel, Fellows; J. J. Frankel, Member; T. J. Higgs, Associate. FINANCE The annual accounts, which follow this report, indicate that income again exceeded expenditure by a margin of R21 410 (last year R28 524). This year's income of R71 184 showed a further increase compared with last year's figure of R55 025. Costs, however, also in¬creased, and it is calculated that, excluding non-recurrent costs, the cost of running the Institute in¬creased by R6 654 from R25 501 to R33 155 for the current year. The main increases were involved in secretarial fees, stationery and print¬ing, and in publication of the Journal. All indications are that costs will continue to rise, and we have been advised that the A.S. & T.S. will be increasing their secretarial fees by 25 per cent in the year ahead to meet their rising costs. This year's surplus was derived entirely from the revenue from sales of publications (R13 644) and a surplus from the 1974 Winter School (RIO 654). There can be no certainty that revenue from these sources can be repeated each year. Our funds are in a satisfactory position because of these worth¬while activities, but the administrat¬ion of the Institute should not de¬pend on them. Council will accord¬ingly have to give consideration in the year ahead to an upward adjust¬ment of subscriptions to meet rising costs.

Jan 10, 1975

By M. A. Jones

Performance of a pilot flood in eastern Kansas indicates improved recovery and accelerated production resulting from mobility ratio control obtained by adding a high moleculur weight polymer to injected water. Ultimate pilot flood recovery of the 75-cp reservoir oil is anticipated to he 350 bbllacre-ft during the six-year flood life. Cumulative recovery to date totals 242 bbl/acre-ft after 30 months of injection. Reservoir parameters and flood performance are disc~rssed in detail to provide a case history relating to the feasibility of polymer flooding. INTRODUCTION The Vernon Polymer flood of the Brazos Oil & Gas Co., located in Sec. 2, 24 S., 16 E., Woodson County, Kans., was begun in late Oct., 1963. Injection of a 500 ppm solution of Pusher* polymer was initiated in a 15-acre pilot to determine feasibility of improving oil recovery by mobility ratio control. Performance to date and the anticipated recovery of 115,000 bbl from the 329 acre-ft pilot indicate efficient displacement of the viscous reservoir oil. Dependence of flood recovery on the displacing fluid-displaced oil mobility ratio has long been recognized. Displacement efficiency of the reservoir oil decreases with increasing mobility ratio. Therefore, economic reduction of this parameter is an attractive approach to improving oil recovery. particularly from viscous oil reservoirs where mobility ratios are often quite high. Waterflood mobility ratios can be controlled by dilute solutions of certain polymers'; which provide a displacing fluid with considerably reduced mobility. Reservoir parameters and flood performance are emphasized in this paper to provide data relating to polymer flood applicability. THEORY The polymer used in the Vernon pilot exhibits the property of unusually high resistance to flow through porous media.'.' A small concentration in injected water effects a mobility reduction considerably beyond that predicted from increases in bulk viscosity as measured in a capillary or rotating cup viscosimeter. Water mobility in the subject flood is reduced by a factor of 7.55 with the addition of 500 ppm Pusher chemical to injected water while the bulk viscosity is increased from 1.03 to only 1.40 cp at 75F. The exact flow mechanism causing resistance to flow has not been established; however, it is observed only in tortuous flow systems and appears to be influenced by brine properties, rock properties and oil saturation. Consequently, mobility reduction is estimated from flow studies for a given sample of reservoir rock and is expressed in terms of a "resistance factor".' This parameter is defined as the ratio of brine mobility to polymer solution mobility measured at residual oil saturation: RF= ?m/?p= (km/µm)/(kp/µp)M.......(1) where RF = resistance factor (dimensionless) k*. and A, = mobility of water and polymer solution, respectively k,,, and k,, = water and polymer permeability, respectively (darcies) µie and µp = viscosity of water and polymer, respectively (poise). Eq. 1 is similar to the expression commonly used to relate the relative mobility of the water bank at residual oil saturation to oil-bank mobility at irreducible water saturation: where M =mobility ratio of water to oil (dimension-less) A, = mobility of oil k, = permeability to oil (darcies) )i* = oil viscosity (poise). From Eqs. 1 and 2 it is apparent that the polymer-solution-to-oil-mobility ratio can be approximated by dividing the brine-to-oil ratio by the resistance factor RF: Mpn=M10/RF...........(3) where M,-, = mobility ratio of polymer solution to oil (dimensionless). This simplified presentation assumes no dilution of the polymer solution by connate water. Also, the presence of a connate water bank separating the polymer solution from the displaced oil is not considered. Displacement of oil by waterflooding is increasingly dominated by viscous forces as the viscosity of the oil displaced increases. When the oil mobility is less than that of the injected water, the oil-water interface is not a piston-like front.3,' Instead, instabilities occur at the flood front in the form of viscous fingers which tend to penetrate the less mobile oil bank. The magnitude of these protrusions, or the degree of channeling, increases

Jan 1, 1967

GENERAL GEOLOGICAL DESCRIPTION The uranium mineral appraisal herein described was made for an ore deposit in New Mexico. The uranium mineralization occurs as roughly tabular to lens-shaped deposits of varying thickness within the sandstone of the Jackpile formation of Jurassic age. The mineralization occurs at several horizons within the Jackpile formation, but is not always confined to specific beds. It often transgresses across stratigraphic layers, as demonstrated when attempting correlation of uranium gamma anomalies between adjacent drill holes. Preferred mineral trends have been established in some parts of the deposit, but are not always predictable when attempting to extrapolate beyond the zones defined by drill hole sampling. MANAGEMENT OF DRILL HOLE SAMPLE DATA All of the drill hole sample data used in the appraisal were obtained from interpretation of gamma-ray logs. A brief description of gamma-ray log sampling is given in Chapter 2. Two preliminary steps are involved in preparing gamma-ray logs into sample data normally acceptable for computer manipulation. 1) The gamma analog trace produced by the downhole probe must be digitized, i.e., counts-per-second (CPS) are recorded for each 0.15 m (1/2-ft) interval across the gamma-ray anomaly. The work is usually done on a "digitizer" while following the analog trace with an electronic stylus on a special tablet or table. The other alternative is to do the work manually. Some of the newest downhole logging equipment produces a magnetic tape output which can be read by a computer. 2) The gamma-ray CPS values must then be converted to equivalent U308 assay values, taking into consideration all appropriate correction factors. The US Atomic Energy Commission, now part of the Department of Energy, developed the theory and algorithm for this conversion in the 1960's, and though the work can be done manually, it is best done with the GAMRED (gamma-reduction) computer program. Fig. 14-1 illustrates a typical analog trace produced by the probe instrumentation, including gamma-ray trace (left), self-potential electric log (center), and resistivity log (right). Fig. 14-2 shows a form of the computer output after the analog data has been digitized to 0.15-m (1h-ft) intervals and converted to percentage U308 values for the same hole. The U308 grade bar graph on the right permits rapid visual inspection of anomalous values by depth and elevation. The converted U308 sample data were next processed by the SURVEY program that assigned to each assay interval [0. 15-m (1h-ft)] an X Y Z coordinate, taking into account all deviations indicated from downhole survey information. The output file of the SURVEY program was listed on the line printer for examination by project geologists, and also became the main input file for the URAN/ COMP compositing program, which is designed to group individual 0.15-m (/-ft) U308 sample values into larger intervals representing discrete anomalous zones according to a series of specified parameters entered into the program by the user. Composite U308 intervals were computed by the URAN/COMP computer program, the logic of which may be briefly described as follows: Step 1 The program searches for the highest U308 value in the hole and builds a composite by adding the highest grade interval on either side so long as the added value is greater than a minimum acceptable grade, and the total composite, after addition, meets minimum GT (grade times thickness) constraints. Step 2 The program loops back to find the next highest assay value in the hole not already included into a composite, until all anomalies above minimum cutoff grade are composited. Step 3 Individual composites are combined if they are closer than a specified minimum waste thickness, and the resultant composite grade still meets GT minimums.

Jan 1, 1980

By P. R. Smith, R. A. Walters

Borates have been mined in the desert areas of California and Nevada for more than 100 years. To about 1890, playa surface mining provided the chief sources of boron minerals. Underground mining of colemanite and later of borax and kernite was predominate until about twenty years ago. Open pit mining of the large deposits of borax and kernite near Boron, California has been most significant for the past twenty years. Mining of colemanite in the Ryan, California area, near Death Valley, began in 1907. Following the discovery of the large deposits in the Boron area (about 1957), mining in the Death Valley area became nearly nonexistent. Only small tonnages were mined for special uses. Little mining was done in the Boraxo area near Ryan. The first claim was made in about 1915. In 1960 the area became the property of the Kern County Land Company, which was acquired by Tenneco Oil Company in 1967. In 1976 the various boiate properties and claims in this region were acquired by American Borate Corporation. The open pit mine is now approximately 122 m (400 ft) deep, 910 m (3000 ft) long and 305 m (1000 ft wide). The borates in the Boraxo pit consist primarily of three minerals. These are about 50% colemanite (CB6011 5H20), about 40% probertite (NaCaB50g 5H20), and 10% ulexite (NaCaBgOg 5H20). The colemanite, along with boric acid and high-grade colemanite ore from Turkey provide the only sodium-free borates for production of textile grade fiberglas. When heated to its decomposition temperature, colemanite decrepitates to a fine powder, which is the basis for the concentration process. The gangue minerals in this deposit are primarily calcite and clays, including bentonite. The ore body has a very low arsenic content, which is a desirable feature. Test work had been done with samples prior to the results discussed herein. This paper will discuss results of test work which were the basis for erection of a plant, and the subsequent plant operations. Laboratory Calcination and Air Tabling Tests Laboratory calcination tests showed that substantial upgrading of the borate could be accomplished by calcining followed by screening of the calcined material. Removal of the + 28 mesh calcine resulted in borate losses of less than 10% with a rejection of 40 weight % or more of the calcine. The minus 65 mesh calcine generally met the requirement of containing 48%, or more, B203. The minus 28 plus 65 mesh material contained an intermediate quantity of borate and would require additional treatment. Testing demonstrated that ore would not have to be reduced to a size finer than 19 mm (3/4 in.) prior to calcination. A temperature range from 400 to 455OC (750 to 850°F) was apparently satisfactory. Calcination at a temperature of 48Z°C (900°F), or higher, was unsatisfactory due to fusion. All laboratory calcination tests were static tests conducted by placing small covered charges in a laboratory furnace for 40 min. In all tests vapor issues from the furnace for 5 to 7 min. Following this period the ore could be heard "popping," due to decrepitation of the colemanite. The reaction generally continued for approximately one-half hour. Various size fractions of the calcination products from laboratory tests were subjected to laboratory air tabling tests, usually after removing the plus 28 mesh material. Laboratory air tabling tests were conducted employing a Whippet V-80 model air table manufactured by Sutton, Steele and Steele Co. now known as Tripple S Dynamics. Variables include both end and side-tilt, speed of vibration, and quantity of air rising through the deck. In addition to the variables in the machine itself, the feed rate is also a rather critical variable. Testing demonstrated that all - 28 mesh size fractions of the calcine could be successfully concentrated to 48% F2O3 or greater. For the finer material recoveries into the concentrate were between 85 and 90% of the borate. With the coarser material a substantial amount of middling was produced which required cleaner tabling. Laboratory calcination and air tabling tests indicated a process whereby the borate could be concentrated to about 50% B203 with borate recoveries approaching 90%. Moreover, the iron content of the concentrate was well below the required specification of 0.3% Fe2O3. Pilot Plant Calcination Following the laboratory test work described above, pilot plant testing was conducted to prove the process, provide data for engineering studies, and provide product for a prospective purchaser. The kiln used was 0.9 m-diam (3 ft) by 9.0 m (30 ft) long and had a belly section 1.2 m-diam (4 ft) by 2.74 m (9 ft) long near the discharge end. The kiln was operated at a speed of 0.7 rpm. Gas was fired into the kiln at an average rate of 27.1 m3/hr (958.4 cu ft per hr). The air to gas ratio used was 10:1. The ore was fed to the kiln countercurrent to the flame and discharged through a hopper into a screw conveyor which discharged to a 1.2 m (48 in.) Sweco separator. The separator had 28, 65, and 150 mesh screen cloths, with the plus 28 mesh fraction being discarded. The minus 28 mesh fractions were later subjected to air tabling. The exit gases, containing some calcine dust, were swept through two cyclones to recover the dust. The gases then were scrubbed in a Ducon scrubber; very little dust reported past the first cyclone. The dust from the first cyclone was also saved in drums. In addition to the gas rate, the flue gas velocity, after

Jan 1, 1981

By S. J. Bluhm

INTRODUCTION Lesotho is a mountainous area of southern Africa from which water is to be exported in an extensive tunnel system, to industrial regions inland. The related tunnelling project has involved a num- ber of drives using tunnel boring machines [TBMs] to excavate about 100 km of 5 m diameter water tunnels [von Glehn and Bluhm, 1995). This paper describes the ventilation and cooling of some of the tunnel drives from both the operational and design points-of-view with a particular focus on heat generation. There were many common features in all of the drives but this paper is focused mainly on the Hlotse drive which was 18,4 km long. The drives were ventilated using forced ventilation systems to provide appropriate air flow throughout the tunnels and face zones. In addition, the Hlotse drive required refrigeration equip- ment which provided chilled water to the tunnel. While the sec- ondary ventilation systems play an important role in gas and dust handling, the paper concentrates on the primary ventilation and cooling issues. The ventilation of these tunnels was an exacting exercise be- cause: • Rock temperatures and geothermal heat flow were high. • TBMs with relatively high power ratings were used. • Diesel locomotives were used. • Drives were relatively long. • High altitude meant a low air density. An important feature was the simulation and monitoring of the ventilation and heat flow components and the project was characterised by analysis, monitoring and ongoing tactical decision-making throughout the progress. The thermodynamics of the systems were complex because there were many interactive effects and analyses were carried out using special computer pro- grams. The monitoring confirmed the accuracy of the models, and in this manner it was possible to confidently ensure healthy and safe working conditions and still minimise costs. Local ambient climate conditions range from temperatures higher than 35 "C in summer to below -10 OC in winter. Based on available statistical data and the thermal storage/damping effects in the system, design summer ambient conditions were taken as 15 OC/25 "C wet-bulb/dry-bulb. The barometric pressure was 80 kPa and due to the altitude, the ambient air density was only 0,9 kg/m3. The local Authority specified a maximum in-tunnel wet- bulb temperature [at any point] of 32,O OC and a mean wet-bulb temperature [from all locations] of 27,5 OC maximum. The maxi- mum height of ground cover above the tunnel was 1 200 m and the maximum virgin rock temperature was 41 OC; see Figure 1. Diesel dilution criteria specified by the local Authority was a minimum of 0.1 m3/s per rated kW of diesel engine. Other requirements related to gases such as CO, CO2, NOx and CH4 [and the need for intrinsically safe equipment] but these are not of direct relevance to this paper. The actual average face advance was about 30 m/d with good days achieving 60 m/d and good months achieving 1 000 m [23 working days]. The original design tunnelling rate was 50 m/d. DESCRIPTION OF HLOTSE DRIVE VENTILATION AND COOLING SYSTEM The ventilation requirements in the tunnels were dictated by heat and diesel dilution needs. The best ventilation and cooling policy is generally a balance between using increased quantities of fresh air or refrigeration [or both]. In this particular scenario it turned out that, since the diesel emission criteria required large quantities of air, the refrigeration needs were modest. The drive was ventilated using a ducted, forced ventilation system from fans located at the portal. The maximum ventilation requirement was 51 m3/s when the drive was at 18.4 km. From a heat flow point of view, the worst scenario was a heat load of 3.5 MW when the drive was at 7 km. This was cooled by the ventilation air and a supply of chilled water to the tunnel. Refrigeration and chilled water system In the design phase, a detailed comparison was carried out between two general alternatives for providing refrigeration. First, was a system in which refrigeration sets and air coolers are installed on the TBM train; the refrigeration sets are cooled by condenser water piped to and from cooling towers at the portal. Second, was a system in which refrigeration water chillers are in- stalled at the portal and chilled water is piped into the tunnel. The detailed comparison indicated that the capital and running costs of the second system were at least 60 % lower than the in-tunnel plant. There were also many obvious practical benefits for favouring the portal system. The refrigeration plant supplied 23 11s of cold water at a temperature of 10 OC. After providing the cooling effect in the drive, the water returned to the portal where it was initially cooled in open-circuit evaporative pre-cooling towers, chilled in the refrigeration plant and then returned to the tunnel. The cold water flowed into the tunnel in an insulated supply pipe and returned in an uninsulated pipe; the water was simply circulated to the end of the pipe and returned. The cooling effect in the tunnel was achieved entirely through heat transfer from the pipe [long linear heat exchanger] and no air coils or other heat exchangers were required. The cooling requirements were satisfied by the heat transfer to the returns chilled water steel pipe [200 mm]. The pipes were eventually installed to a maximum distance of 10,8 km in what was considered a very practical and cost effective solution.

Jan 1, 1997

By Rolf Pollak, Jan Tegtrneier, Gerhard Keller, Horst Gerhardt, Wolfgang Müller, Dieter Tetzner

INTRODUCTION Under certain geological and mining conditions mining near the surface results in an intensive destruction of the overlying rock strata. If the system of mine openings remains filled with air after having completed all mining activities because the flooding of the mines depending on the morphology of the mining area does not exceed a certain level, this circumstance might result in dangers for the public. This is particularly true where uncontrollable air- ways occur within the destroyed rock or if former uranium mining areas are concerned. In these areas radon emanation resulting in undue radon concentration should be taken into account. This circumstance can be counteracted by suitable measures of ventilation. In the following the example of a uranium deposit in the Erzgebirge Mountains (Saxon Ore Mountains) is used to illustrate how statements about the ground flow through overlying rock can be made on the basis of systematic investigations. Finally comparisons with experience gained in coal mining yield orientation data which are used as input quantities in ventilation network calculations and in project preparation of mine fans. Last not least it is proved that endangering the population by radon can be excluded by installing an appropriate fan. INITIAL SITUATION From 1946 to 1989 the uranium deposit of Schlemal-Alberode was mined starting from superficial zones down to depths of 1,800 m. The total output amounted to 80,000 t of uranium, i.e. Schlemal Alberode is one of the largest uranium deposits in the world. The mine openings comprise a cavity of 40 million m3. Al- though thin seams (0.2 to 0.3 m) and steep to semi-steep veins (60 - 70 degrees on average) were extracted using the method of overhand cut-and-fill there were also deposit sections with larger thicknesses and poor ore zones, respectively, where especially after the war until about 1955 large openings were formed due to missing filling. Due to this uncontrolled, relatively "wild" mining considerable subsidence of the surface and fractures in the border zone of the subsidence area occurred. The town of Oberschlema was destroyed and an area of 0.14 km2 was damaged by mining involving a maximum subsidence of 6 m in the centre - called deformation area. The rehabilitation work for the deposit area of Schlemal Alberode involves the flooding of the openings up to a level of the Markus-Semmler-Sohle as long-term preventive measure. Fig. 1 shows the flooding level of October 1996. The level of the -540 m drift will probably be reached in June 1997. From that point of time a general rearrangement of the ventilation has to be made because the central return air ventilation via shaft 373 will no longer be possible. This shaft has no connection with the superficia1 system of mine openings. A new return air ventilation system based on shaft 382 as return air shaft will be designed (Fig. 1). The capacity of the future mine fan system will e.g. be determined by the flow conditions in the deformation area mentioned above. For this reason investigations were carried out in a test drift 13a on the level of the Markus-Semmler-Sohle in the stretching zone of the subsidence trough for several days (Fig. 2 and 3). INFLUENCE OF THE OPERATING CONDITIONS OF THE MAIN MINE FAN ON THE RADIATION SITUATION ON SURFACE The beginning of flooding in 1991 was also the beginning of the investigations into the options of future ventilation due to the reduction of the mining area. The first analyses of radon emanation made in May 1991 showed that only about 20 % of the total radon emanation result from systematically ventilated mine workings. This means that despite of existing dams the air streams through the majority of mine openings of the deposit. The analysis of the individual radon inflows illustrates the special problematic nature of the deposit involving all possible effects on the surface. The shutdown of the main mine fan over the long period between 2.10. and 6.10.1991 showed the essential influence of the operating conditions of the fans on the whole radiation situation on surface. Fig. 4 shows the comparison of the radon concentration level and its daughters during a 'normal" weekend with the level during long-term shutdown in the old 'Casino" of the papermill of Niederschlema. This papermill as well as the 'Casino" are situated in the outcropping zone of the deposit section of Niederschlema. The -60 m level was driven as the upper level in this area, from which the panels were driven to about 20 m below the surface. Also below the "Casino" there is such a panel in the vein 'Glirck", whose roof is directly situated below the building in a depth of 39 m. Directly next to it there is the raise 2 of the panel 'Namenlos" in a depth of 27 m. The graphics clearly shows the immediate increase in radon concentration as well as in the daughter concentration after the shutdown of the fans. The elimination of the artificially generated negative pressure results in the reversal of the convection direction due to the natural ventilating pressure. The interface structure allows the mine airto pass through aero- dynamic connections to the surface. Neglecting the meteorological parameters the concentrations of radon and its daughters in- creased by about 350 % as compared to a 'normal" weekend during longterm shutdown (i.e. in 4 days). Furthermore it could be noticed that a stationary state between the radon and its daughters has not been established in the period

Jan 1, 1997

The purpose of this paper is to outline some aspects of the experimental worlr that has been carried out in the mineral dressing laboratory at M'ount. Morgan during recent years.The subject matter has been dealt with under the following headings:1. Comminution of various mill products.2. The application of sulphur dioxide as a flotation reagent.3. Heavy-media separation.4. Bulk flotation of sulphide ores.COMMINUTION OF VARIOUS MILL PRODUCTSThe following data were cbtained by processing fractions of reserve grade ore.Granular TailingIn Table 1 are shown the results of tests on material representative of average grade tailing. In the first test (H1 ) the granular tailing was floated as received and in the second test (Ha) the tailing was ground prior to flotation.

Jan 1, 1952

By Kelly J. Driscoll

Impure zinc (98.5%) was electrorefined to high purity zinc (99.995%) using a fused salt electrolyte of molar composition 1: 1 : 1 :2 ZnCI2:KCl:NaCl:LiCl with a current efficiency of about 90% at 723 K. A recessed channel electrorefining cell was developed which achieved excellent mixing of both the metal and the electrolyte and by ensuring a very small anode to cathode spacing the voltage drop across the cell was 0.16 V and at a current density of 7kA m-2. The physical characteristics of the cell were observed using a low temperature model which showed the formation of droplets and rivulets which provided mixing. The application of this approach to the refining of other non-ferrous metals is discussed.

Jan 1, 1992

By Donna D. Harbuck

The Bureau of Mines has investigated a process to produce titanium powder directly from titanium salts. If successful, such a process could lower the cost and increase the use of titanium. In this study, K2TiF6, K2TiF3, and TiF3 were reduced with potassium in the temperature range of 650° to 750° C. Under certain conditions, titanium powder 1.6 to 3.3 microns in size was produced. After leaching the product with a 0.1M HCl solution, powder impurities as low as 0.04 pct potassium and 0.28 pct fluorine were obtained. However, oxygen levels of over 1 pct and the frequent formation of insoluble byproducts demonstrate the need for further research to make the powder suitable for powder metallurgy applications.

Jan 1, 2011

By David C. Lynch

FeCl3 particulate of 10 to 100 um in diameter was reacted with radicals produced in an oxygen plasma at pressures between 1 to 20 Torr. The particulate, in a fluidized bed, was positioned above the plasma to avoid over heating the oxide product which is known to lead to crust and clinker formation[ 1 ]. Reaction between the iron chloride and the gas phase occurred at temperatures between 66 and 226 °C. Oxygen radicals react with the surface material to a depth of approximately 2.5 4m. Surface material, analyzed with a x-ray technique, indicates that the maximum amount of FeCl3 reacted on a weight basis was found to be approximately 67%. More typically, 40 to 50 wt% of the chloride was reacted within 30 minutes to 1 hour.

Jan 1, 1996

By B. Yarar, J. Kaoma

[Introduction The critical surface tension of wetting of hydrophobic materials has been investigated extensively by Zisman et al. (1973) and relates the spreading of a liquid on a solid to the surface tension of this liquid. Thus, the Zisman ap¬proach shows that a plot of Cos 0 versus ?LV where, 0 is the contact angle and ?LV is the surface tension of that liquid in contact with its own vapor at equilibrium, produces a straight line which can be extrapolated to Cos 0 = 1. The intercept corresponds to a given ?LV, termed the critical surface tension of wetting. In instances, however, Cos 0 - ?LV plots show deviations from linearity. Alternative approaches have also included the plotting of adhesion tension (YLV COs 0) versus ?LV that can be interpreted in terms of critical surface tension data (Lucassen-Reynders, 1963). It has, on the other hand, been customary to correlate contact angle (0) with flotation data, although it is recognized that surface inhomogeneities and related contact angle hysteresis limit the interpretations pertaining to solid surface structures purely based on contact angle measurements. It is nonetheless true that 0 = 0, when the liquid is water, corresponds to a hydrophilic solid. Determination of ?c by contact angle measurements in conformity with the Zisman approach can encounter experimental difficulties, especially when solids are in a powder form. Following the reasoning that 0 = 0 corresponds to the hydrophilic state of a solid, froth flotation experiments can be expected to produce critical surface tension of wetting values for powders. The technique that consists of plotting recovery (% R) versus YLV data and extrapolating the linear part of the curve to % R = 0, has been described elsewhere (Yarar and Kaoma, 1984). This paper describes some of the results obtained by the extrapolation technique mentioned above, using various sulfide minerals. Experimental Materials Molybdenite (MoS2), sphalerite (ZnS), chalcopyrite (CuFeS2), chalcocite (Cu2S), and galena (PbS) were handsorted from massive samples and ground by mortar and pestle, and the -147 Fan fraction used in experiments; sulfur (S) was a C.P. Fisher Scientific Co. product. Method One gram powder samples were conditioned for three minutes in methanol-water solutions of predetermined surface tension values and the percents recovered by flotation using a Partridge-Smith Cell (Partridge and Smith, 1972), were plotted against each ?LV value. Results and Discussion The percent recovery versus ?LV for two naturally hydrophobic solids, i.e., natural molybdenite, MoS2 and sulfur are given in Fig. 1 where it is seen that ?c for sulfur is obtained as 31.5 mNm-1 whereas for MoS2, it is 26 mNm-1. Figure 1 also shows that the new flotation technique described, for estimating ?c is in good agreement with the Zisman method. Curves obtained by the same method for the minerals Cu2S, PbS, and ZnS are given in Fig. 2, where it is seen that for Cu2S, ZnS, and PbS, the extrapolation to % R = 0 produces ?c values of 36, 39, and 49 mNm-1, respectively. CuFeS2, on the other hand, gives a curve that does not permit such an extrapolation to obtain a ?c value with certainty. Nonetheless, treatment of the powder with 8.95 X 10-5 molar Na2S solution followed by washing with distilled water prior to flotation effects the curves as shown in the insert of Fig. 2. The type of curves given in Fig. 1 appear to be typical of hydrophobic solids as can also be observed in Fig. 3. The composition of solid and minute quantities of inclusions seem to play an important role as regards the surface properties of solids as reflected by the slope of the linear part of the % R - ?LV plot and thus the value of ?c obtained by this technique. For example, MoS2 of three different stoichiometric compositions produced the data given below. [MoS2MoS2.14Mo1.002S2 ?c mNm-126.029.031.0 dR/d?8.95.23.0]]

Jan 1, 1985

By R. E. Le Messurier

GENERAL DESCRIPTION The mines are located at Broken Hill in the western part of New South Wales, Australia, 930 km west¬northwest of Sydney. Mining Methods Where pillars must be left between stopes, extraction of a block of ore between two or more levels is carried out in stages: 1) Stopes are mined overhand by horizontal cut¬and-fill stoping, or by long-hole stoping. The various methods of horizontal cut-and-fill stoping used are desig¬nated as: -Loader cut-and-fill stoping with horizontal blast¬holes and diesel load-haul-dump (LHD) units. -Open stoping with horizontal blastholes and com¬pressed air or electric scrapers, or with diesel or electric LHD units. Square-set timber stoping, where ground condition precludes open stoping. The backs are sup¬ported with timber square sets. Pattern rockbolting of the backs may be used as an intermediate stage between open stoping and square-set stoping. 2) Vertical pillars are mined by undercut-and-fill stoping or by vertical crater retreat stoping. 3) Where a horizontal pillar has been formed below the level above the primary stope, this level pillar, usually 6.1 to 8.2 m in depth, may be mined by over¬hand square-set stoping or by undercut-and-fill. Long-hole stoping is used at the Zinc Corp. (ZC) and New Broken Hill Consolidated (NBHC) to mine parts of the zinc-rich A and B lodes. Drilling is carried out from drill drives and, in some cases, sills. Hole sizes used are 165 mm (61/2 in.), 110 mm (4'h in.), and 65 mm (21/2 in.) nominal diam with hole lengths being limited to 50, 30, and 20 m respectively. Ammonium nitrate-fuel oil (ANFO) and water gel explosives are used. Broken ore is loaded from drawpoints at the bottom of the stope using rubber-tired diesel-driven LHD units which deliver ore to an orepass. Annual production at Zinc Corp. and NBHC accord¬ing to mining method is given in Table 1. At Zinc Corp. mining is being carried out at depths varying between 200 and 832 m. At NBHC it ranges from 528 to 915 m below the surface. The average travel time to the stope is 30 to 45 min. GENERAL ORE BODY REQUIREMENTS Size, Shape, and Dip There are six separate lodes which carry minable ore in the mines of Zinc Corp. and New Broken Hill Consolidated. The lowest of these and the main ore bodies presently being mined at Zinc Corp. are the lead lodes, consisting of a siliceous fluoritic ore body (No. 3 lens) and overlying it a calcite rhodonitic ore body (No. 2 lens). Each possesses characteristic ratios of lead, silver, and zinc, although the grade within each lens is variable, in places being as high as 30% Pb. There are also low¬grade portions which consist chiefly of poorly mineral¬ized rhodonitic material. No. 3 lens terminates within the Zinc Corp. leases, while No. 2 lens continues into the NBHC leases where it continues to provide a significant proportion of pro¬duction from the mine. Overlying the No. 2 lens are the various zinc-rich ore bodies. In ascending sequence these are lower No. 1 lens, upper No. 1 lens, A lode, and B lode. A typical assay for the small No. 1 lens group ore bodies is 8% lead, 50 g/t silver, and 20% zinc. The `A' lode is a large low-grade ore body in which the higher grade sections assay 4% lead, 30 g/t silver, and 10% zinc; larger quantities of lower grade ma¬terial also occur. B lode, the main zinc lode, typically assays 5% lead, 30 g/t silver, and 17% zinc. B lode increases in size as it extends south, and in NBHC is a major ore body, contributing about 52% of the production. All the lodes pitch south, the average pitch being 0.52 rad (30°) in the Zinc Corp. leases and rather flatter than this in the NBHC leases (see Figs. 1-3). The ground is all competent rock. Production requirements at Zinc Corp. are 900 000 t and 1 200 000 at NBHC.

Jan 1, 1982