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By R Uken G P Whitmore

The coastal dune cordon of northern KwaZulu-Natal is amongst the highest in the world reaching heights in excess of 180 m and up to 2 km in width. It comprises a complex system of Quaternary stacked dunes of various geomorphologies, types and ages. The dunes contain economic heavy minerals which are currently exploited ~20 km north of Richards Bay. Although mining of these ore sands has been in operation for the last two decades, little work has been undertken on the clay-fraction (

Jan 1, 2001

By Ehsan Vahidi, Fereshteh Rashchi, Amirhossein Shahnazi

"The pig iron from steel manufacturing plants contains a high concentration of vanadium-bearing materials smelted with iron ore. During the oxidation process of the molten pig iron with oxygen lances, vanadium transfers to the slag. In this research, recovery of vanadium from LD (Linz-Donawitz) converter slag of steelmaking plant was investigated. The leaching residue was characterized by XRD and XRF techniques. The maximum vanadium recovery was achieved at optimum leaching conditions of 70 °C, S/L: 1/15, sodium hydroxide concentration: 3 M and leaching time: 150 minutes. The dissolution rate increased with rising sodium hydroxide concentration and temperature and decreasing particle size. The experimental data are treated graphically to explain the kinetics of the vanadium recovery process using shrinking core model (SCM). As a result, the controlling regimes in the SCM were analyzed separately using liquid film diffusion control, solid product diffusion control, and reaction control mechanisms.IntroductionVanadium is commercially important as a constituent of several alloys and catalysts. Nowadays, vanadium is mostly recovered as a by-product from secondary sources or from industrial waste streams such as titaniferrous magnetite, fly ash, spent catalysts, and petroleum coke. Steelmaking operations are specifically concerned by this problem due to the production of a huge quantity of by-products [1]. These by-products, including LD converter slag, sludge, fly ash, spent catalyst, ash, and alloyed scrap, contain a notable amount of heavy metals which may cause environmental problems. Solvent extraction is one of the techniques being increasingly used for the recovery of vanadium [2]. Apart from vanadium ore, secondary resources such as titanmagnetite, ilmenite and slag from the ferrous industry are a major source of supply. The pig iron from steel manufacturing contains a majority of vanadium-bearing materials smelted with iron ore. In a heat resistant shaking ladle the molten pig iron is oxidized with oxygen lances, causing the vanadium to be transferred to the slag. Normally, the vanadium in LD converter steel slag (coming from Linz-Donawitz steelmaking process) is ca. 5% as vanadium pentoxide (V2O5). This slag is the world's principal raw material for vanadium production [3, 4]. Moskalyk and Alfantazi [4] reported that a roast-leach process is used in South Africa for the production of vanadium oxides from magnetite ores while some of the slag was converted to ferrovanadium. A portion of the slag was exported to Europe for the production of vanadium oxide. The researchers mentioned that alkaline roasting was the main process at that time employed to extract vanadium pentoxide from vanadium slags."

Jan 1, 2012

By Jian Li, S. Penttilä, P. Liu, W. Zheng, C. Bibby

"The materials selection for Gen IV supercritical water reactor in-core components faces major challenges due to severe operating condition. The current Canadian pressure-tube supercritical water (SCW) concept design requires core outlet temperature of 650 °C at 25 MPa with peak cladding temperature reaching as high as 850 °C. Aside from oilier challenges, there is currently very limited data on candidate stainless steels' corrosion resistance under this operating condition. Austenite stainless steels that include 316L, 304H, AL6XN and 3 l0S that cover a wide range of Cr contents are promising candidates for either direct in core applications or as base coating materials. However, their corrosion resistance has not been tested systematically under this stringent condition. In this study, these types of stainless steels are tested, and their corrosion resistances are compared.IntroductionThe Canadian pressure-tube supercritical water (SCW) concept will operate at a core outlet temperature of 650 °C at a pressure of 25 MPa, with potential peak cladding temperature reaching as high as 850 °C. High-temperature corrosion resistance is one of the key design requirements for fuel cladding components. As summarized by Allen [1 ], most testing to date has been carried out at temperatures below 730 °C. To date, SCW corrosion data above 600 °C is very limited. Even at relatively low temperature exposure to SCW (e.g. 500 °C), F/M steels develop a thick oxide layer rapidly, while some austenitic stainless steels develop thinner surface oxide. Ni based alloys appear to form very thin surface oxide, however pitting was observed in the vicinity of intermetallic precipitates [2-5]. Dissolution of major alloying components from all of these alloys could also pose significant problems for downstream piping [6]. Kaneda et al. [7] suggested some modified stainless steels provide better SCW general corrosion resistance, with no sign of SCC under the slow-strain rate test.Cho [8] suggested that as the Cr content increase, general corrosion resistance improved at 510 °C and 25 MPa. However, similar to Cho's work, most of the SCW corrosion resistance assessments were based on weight change measurement. This can be unreliable due to potential exfoliation of thick oxide film. Besides, there have been no systematic SCW corrosion resistance tests for candidate stainless steels under higher temperatures. SCW corrosion rate increase dramatically for alloys tested at 600 °C [1]. It is also important to understand the role of major alloying elements under this stringent operating condition, not only for fuel cladding components' materials selection, but also for coating materials development."

Jan 1, 2013

By Randolph E. Scheffel, David C. Fitch, David K. Hogan, Mark R. Welsh

In Oct. 1981, Ranchers Exploration and Development Corp. began producing silver bullion from the Escalante mine and mill complex, located in southwestern Utah. Current in-place reserves are estimated at 2.42 Mt (2.67 million st) of ore, containing about 936 t (33 million oz) of silver at an average grade of 428.5 g/t (12.5 oz/st). The vein-type deposit is being mined underground through a decline that is currently 864 m (2,835 ft) in length, but will ultimately be at least 1.5 km (5,100 ft) long. The mine is completely mechanized with rubber-tired equipment, and is producing 680 t/d (750 stpd) of ore, five days a week, or 177 kt/a (195,000 stpy). Additional capacity is available. The cyanide leach mill and associated refinery are presently producing 4.35 t/month (140,000 oz/month) of silver bullion-an an¬nual rate of about 52.25 t (1.68 million oz). With minor modifications, the mill can produce up to about 70 t/a (2.25 million oz/year), and expansion is currently underway. History The Escalante property is located 68 km (42 miles) west of Cedar City, UT, on the west side of the Escalante Valley, where the vein forms a gentle ridge in an area of low, rolling hills. For years, the vein has been the object of considerable exploration, but never became more than a sporadic producer, primarily due to low silver prices. Heber Holt of Enterprise, UT, staked the original four claims in 1896; they were patented in 1910. During the next 48 years, limited exploration and mining activity was conducted on the property. Systematic evaluation and development did not, however, begin until March 1958, when Samuel S. Arentz, a prominent mining engineer and consultant, formed several joint ventures with mining and oil companies that conducted drilling, drifting, mining, and water drawdown tests on the property. However, it remained uneconomic, principally because of low silver prices and costs for pumping large flows of underground water. In 1975, Ranchers obtained an option on the property and in 1979 began a drilling program. As of May 1982, the company had drilled 70 core holes, bringing the number of exploratory holes drilled on the property to 138. Test mining began in 1979 as silver prices increased, and full mine development began in 1980. Mill construction began in Sept. 1980, and on Aug 29, 1981, the first ore was fed to the mill.

Jan 9, 1982

By H. Nemutolluhi

Discussion by G.E. Agar In the April issue of Mining Engineering there is a technical note "New size laboratory ball mill for Bond Work Index determination" written by M. Nematollahi which is very confusing. My recollection is that a description of an experimental procedure should be such that someone skilled in the art should be able to reproduce the results. Applying this test to this particular article would according to me cause a failure to be registered. There are apparently three batch laboratory test procedures referred to in the article. One of these is the traditional Bond test with all its limitations such as dry grinding followed by exclusively dry screening, two attributes which fail to simulate conventional industrial grinding. There is a second procedure (NSBM) which the author is advocating as an improvement of the Bond procedure but from the description I doubt that anyone would be able to repeat the procedure. For instance, is the grinding to be done wet or dry, if wet at what pulp density, and how is the screening to he done, wet or dry or a combination? On top of that uncertainty the author refers to a simplified "Anaconda" test. No reference is provided for this procedure and in the text the results of the technique are referred to "Denver wet ball mill" results and say nothing of Anaconda results. Reply by H. Nematollahi I think that the method proposed to determine the Bond Work Index in this paper is clearly described in the paper as follows: •The method of grinding is dry (Ref. paragraph I, subject Method description); •The size of ball mill, the quantity and size of balls are noted (Ref. paragraph 1, subject Test equipment); and The test procedure is described in detail (Ref. paragraphs 1, 2 and 3, subject Test procedure). As it is noted in the paper, the test has been worked out under the same condition as Standard Bond Test with ball mill, so dry (screening should be dry also), but using a smaller ball mill and a smaller quantity of sample. Considering the above mentioned points, I believe that the method has been fully and clearly described and it ought to be repeatable as well. The Anaconda test is a method for determination of Work Index and could be done by any types of laboratory wet ball mill (Yan, R.F., et al., cf. References). I applied this method using a laboratory Denver ball mill (Ref. paragraph I subject Determining the Bond Work Index). So the results referred to Denver wet ball mill can be equated the same as Anaconda method.

Jan 1, 1996

OFS Branch Proceedings of the Annual General Meeting ofthe O.F.S. Branch on Wednesday, 31st July, 1974. The Annual General Meeting of the Branch was held at the Western Holdings Limited Canteen on Wednesday, 31st July, 1974 at 7.30 p.m. Mr E. T. Wilson (Chairman) was in the Chair. There were also present: Papers of interest The following papers may be of interest to members. Education fund The Institute has recently brought into being a Fund to finance various educational projects.

Jan 5, 1974

An investigation into the stability of large open stopes at Premier diamond mine, Culinan, South Africa, is described. The open stopes are located in a weak kimberlite orebody, directly underneath a 75 m-thick sill of gabbro. Two and three-dimensional stress analyses were carried out in a back-analysis of the stresses and failure around the first open stope created under the sill. The results were used in the prediction of the future stability of the planned open stopes, intervening pillars, and sill.

Jan 1, 1987

Coal Division Mineral Industry Education Division Industrial Minerals Division Minerals Beneficiation Division Mineral Economics Division Mining, Geology and Geophysics Division Council R C Stephenson, Chairman (Ind Min Div), M D Cooper (Coal), H F Yancey (Coal), S D Michaelson (MBD), Will Mitchell, Jr (MBD), J R Van Pelt, Jr (MIED), S S Cole (Ind Min Div), J D Forrester (MGGD), L E Shaffer (MIED), E L Clark (MGGD), C W Merrill (MED), E H Robie, Acting Secretary, 29 West 39 Street, New York 18, New York

Jan 1, 1955

By David Green

"This paper describes Outokumpu's Tank Cell concept using the unit reactor principle and operating experience gained in various applications. Presently the tank cells used range in size from 35 t pilot size to 100 m' (3500 cuft). The paper discusses operating experience and results from commercial installations in copper and zinc rougher and cleaner services.The key feature of the Tank Cell described is the effective control of the froth phase and resulting possibility to operate on an improved grade recovery curve by manipulating the control variables."

Jan 1, 1995

New York Holds monthly meetings, except June, July and August JOHN A CHURCH, JR, Chairman JOSEPH E POGUE, Vice-chairman SIDNEY ROLLF, Vice-chairman EDWARD C MEAGHER Secretary, Texas Gulf Sulphur Co, 41 E 42nd St, New York M H MERRISS, Treasurer, 42 Broadway, New York ARTHUR NOTMAN T T READ Boston Meets first Monday of each winter month EDWARD C GEITHER, Chairman WILLIAMS BLACK, Vice-chairman DONALD H MCLAUGHLIN, Secretary-treasurer, Rotch Bldg, Harvard Univ, Cambridge, Mass J L GILLSON A G WAITD

Jan 1, 1923

By P. Campbell, E. M. Gardy, L. Hachkowski

Introduction The efficient process control of a uranium mill and its associated mining operation requires a nearly continuous knowledge of the uranium concentration in the ore. Generally, the approach is to use laboratory assays of grab samples from the mill feed belt. In some cases, elaborate and expensive systems have been used to ensure random sampling, but even with these systems, mass-balance discrepancies still exist. There is a requirement for an on-line instrument that can give a prompt, accurate analysis of a large portion of the feed stock. The authors have recently evaluated a laboratory system that achieves these goals using neutron activation and delayed neutron counting. The development of the on-line uranium ore analyzer is a consequence of previous work done at the Whiteshell Nuclear Research Establishment (WNRE) (Campbell et al., 1978 and 1981), and is based on the emission of delayed neutrons after the neutron irradiation of fissionable materials. The mechanism of delayed neutron emission has been described (Keepin, 1965), but briefly is as follows. The fission fragments resulting from irradiation are in an excited state. Certain of these delayed fission products, precursors, decay to a more stable state by the emission of a delayed neutron. The delayed neutrons can be divided into six groups with effective half-lives ranging from 0.2-55 sec. It is important to note that more than 50% of the delayed neutrons are emitted within the first 6 sec after irradiation; this has an influence on the design of the analysis instrument. Measurement Cycle The measurement cycle involves a neutron irradiation, a delay time, and a counting period. The laboratory demonstration model of the ore analyzer comprises a californium-252 neutron source, a source chamber, a pneumatic transfer facility, and a neutron irradiation and counting module (Fig. 1). The ore sample is placed in the irradiation chamber. The neutron source is transferred to the center of the chamber for a 5-sec irradiation while the helium-3 thermal neutron detectors are switched off. As the source returns to the storage chamber, the detectors are switched on and the delayed neutrons counted. The total assay cycle is 16 sec. The current instrument design calls for the ore to be cut from the main belt every 16 sec by a mechanical arm and dropped into the analyzer directly, or via another belt. The cycle starts when there is a sufficient mass of ore in the irradiation chamber to depress a switch and stops when the delayed neutron count is finished. The ore is then dropped through the bottom of the analyzer chamber and returned to the main belt. The counts from each cycle are accumulated and the data reduced to give the desired information. Key for Fig. 1 A-Helium-3 neutron detector B-Wax moderator for neutrons C-Irradiation chamber D-Funnel for receiving ore E-Gate F-Transducers G-Biological shield for neutron source H-Transfer tube for neutron source I-Neutron source J- Microprocessor K-Accumulator for hydraulics L-Control panel M-Electro-hydraulic servo valve to control movement of the source N-Electro-hydraulic proportional valve to control movement of the gate O-Hydraulic power unit Irradiation and Counting Equipment Irradiation of the ore is achieved using a 1 X 10 7 neutron*s-1 californium-252 neutron source. The source is contained in an iron capsule and pneumatically transferred between the irradiation/counting and storage chambers. The irradiation/counting chamber possesses 12, 600-mm (24-in.) long helium-3 thermal neutron detectors embedded in an annulus of wax (Fig. 1). Two sets of six detectors are each interfaced to a preamplifier, amplifier and single-channel pulseheight analyzer system, and the two systems are connected to a scaler. A special timing unit controls the transfer of the source and the on and off switching of the detector assembly, in sequence with the movement of the source. Results and Discussion In the experimental demonstration, using the 107 neutron.s-1 source, a 10% uranium ore was measured to ± 10% (2o total error) in 10 cycles and ±6% (2o) in 100 cycles. Since a larger source was not available, it can be statistically predicted that a 109 neutron-s-1 source could be used to measure a 0.1% ore to ± 10% (20) in 10 cycles. The uranium can occur as discrete nonhomogeneous particles in the ore, and the error caused by such a distribution of particles is ± 4% (2o), as determined by

Jan 1, 1984

Discussion of the previous paper E. N. H. MOLLOY The first 22-foot-diameter Aerofall mill was installed at Mangula in 1957. This was fitted with cast-iron trunnion'! located to the mill shell with 2-inch-diameter mild-steel bolts. O.F .S. Branch Minutes of the General Meeting held at the Welkom Club on Wednesday, 2nd June, 1976, at 4.00 p.m. Mr G. J. C. Young (Chairman of the O.F.S. Branch) was in the Chair. Also present were:

Jan 9, 1976

The papers that appear on pages 326 through 367 were presented at two joint meetings of the Mineral Industry Education Division, Geophysics Committee, the American Geophysical Union and the Committee on College Curricula of the American Association of Petroleum Geologists, at the A.I.M.E. meeting in New York on Feb. 17 and 18, 1941. At the first meeting, the Mining Geology Committee, the Society of Economic Geologists and the Committee on Applications of Geology of the ilmerican Association of Petroleum Geologists also took part.

Jan 1, 1946

By K. A. G. Blendulf

The paper describes tests on horizontal-belt filters for the filtration of gold and uranium. The promising results led to the installation of 17 such filters (ten of them 120 m2 in size) in the mine's metallurgical plants, and their operation is discussed. Although several problems were encountered both in operation and maintenance, it is concluded that, with correct operation and suitable filter cloths, exceptionally good metallurgical recoveries can be achieved at filtration rates twice to three times higher than those on rotary filters.

Jan 1, 1997

By J. K. E. Douglas

Brigadier Ralph Shelton Griffin Stokes D.S.O., C.B.E., M.C., who died in England on 24th February, 1979, at the age of 96 years, was one of the most colourful personalities in South African mining history. After the South African War, in which he served as a trooper, he joined the Crown Reef Mine as an assistant assayer (1902), and was to serve the mines of the Corner House Group intermittently for the next 57 years until his retirement in 1959.

Jan 1, 1979

The papers that appear on pages 326 through 367 were presented at two joint meetings of the Mineral Industry Education Division, Geophysics Committee, the American Geophysical Union and the Committee on College Curricula of the American Association of Petroleum Geologists, at the A.I.M.E. meeting in New York on Feb. 17 and 18, 1941. At the first meeting, the Mining Geology Committee, the Society of Economic Geologists and the Committee on Applications of Geology of the ilmerican Association of Petroleum Geologists also took part.

Jan 1, 1946

CENTRAL AMERICANIZATION COMMITTEE Chairman, MRS. C. C. BURGER This committee reports progress in drawing up circulars on the. subject of Americanization, to be sent to all the branches of the W. A. A. I. M. E. It is greatly to be desired that all our members should interest themselves personally in work that can be done among the women and children in the mining districts in the way of Americanizing those of foreign birth or extraction; so much can be accomplished by individual interest and sympathy.

Jan 11, 1918

By D. W. Lillie

It has been demonstrated that considerable bend ductility exists in bulk specimens of polycrystalline high-purity silicon. The possibility of hot-forming at 1200°C is suggested. EXCELLENT corrosion resistance in many media and low cross section for absorption of thermal neutrons (0.13 barn) would make silicon of interest to nuclear engineers were it not for extreme brittle-ness and the difficulty of fabrication by any reasonable means. The use of silicon for structural purposes also has been considered in view of its light weight and oxidation resistance. Johnson and Han-sen' have investigated the properties of silicon-base alloys and concluded that there was no way of making pure silicon or silicon-rich alloys ductile at room temperature. In view of reports of appreciable ductility in germanium single crystals above 550°C'." and some plastic deformation in single-crystal silicon above 900oC,' the present investigation was undertaken to define more precisely the limits of high-temperature ductility in pure silicon. After this investigation was begun torsion ductility in both germanium and silicon was reported by Greiner." Through the courtesy of F. H. Horn, a small bar of cast extra high-purity silicon was obtained and small bend specimens were made from it by careful machining and grinding. All of the reported tests results were obtained from samples from this bar (bar No. 1) and one other of similar source (bar No. 2). No complete analysis was obtained but, based on analysis of similar semi-conductor grade material, metallic impurities were under 0.01 pct total. Vacuum-fusion analysis for oxygen showed a value of 0.0018 2 0.0003 pct for the first bar tested and metallographic analysis showed no evidence of a second phase. Bend tests were carried out on an Instron tensile machine using a bend fixture with a 1 -in. span loaded at the center. Supporting and loading bars were 0.250 in. round and the load was applied by downward motion of the pulling crosshead of the machine. Specimen thickness and width were approximately 0.10 in. and % in. respectively. Loading rate was controlled by holding crosshead motion constant at 0.02 ipm. In some cases a smaller specimen was used on a 5/8-in. span with a 0.129-in.-diam loading bar. The entire bend fixture was surrounded by a hinged furnace and all heating was done in air atmosphere. Temperature measurement was made with thermocouples fastened directly to the bend fixture within less than 1 in. from the specimen. Autographic stress-strain curves were recorded during each test, and breaking load, total deflection, and plastic strain could be obtained from these curves. Stress was calculated from the beam formula S = 3PL/2bh2, where P is the load in pounds, L the span in inches, b the specimen width in inches, and h the specimen thickness in inches. This formula is strictly correct only in the elastic range but has been used to calculate a nominal stress for convenience in the plastic range. The stress given is the maximum stress in the specimen. Results The results of the complete series of tests are shown in Table I. The first group of tests (specimens Nos. 1-6) showed the beginning of plastic flow at a test temperature of 900°C, so two additional tests (Nos. 8 and 9) were made at 950°C on small-size specimens from bar No. 2. Specimen No. 8 was tested in the as-machined condition, and No. 9 was heat-treated in hydrogen at 1300°C for 2 hr, cooled to 1200°C and held 1 hr, cooled to 1000°C and held 1 hr, cooled to 900°C and held 1 hr, and finally cooled to a low temperature before removal from the hydrogen. It is apparent that the heat-treatment had a significant effect on yield strength and ductility. In addition, the magnitude of the yield point was conslderably reduced in the heat-treated specimen as is shown m Fig. 1 by tracings of the stress-strain curves. After obtaining a furnace capable of reaching higher temperatures specimens Nos. 10 to 13 were tested at 1100 and 1200°C. Strain rate was increased by up to a factor of 10 to see whether the ductility observed was excessively strain sensitive. Specimen NO. 10, strained at 0.02 ipm and 1100oC, was still bending at a deflection of 0.322 in. when the load rate was increased to 0.2 ipm, resulting in immediate

Jan 1, 1959

By Norman Lamont

The evaluation of certain reser-voir properties, such as porosity and fluid saturation, from electrical well surveys has been widely accepted in petroleum engineering. Various investigators have established relationships between these properties and certain parameters which affect the response of the electrical log. Among these are the resistivities of the mud, its filtrate, and its filter cake. In 1949, Patnode1 established a relationship between the resistivities of the mud and filtrate. The well logging service companies have contributed relationships for the mud-mud cake resistivities2,3 These have been valuable since it was the practice to measure only resistivity of mud at the well site. During the mid-1940's the industry began drilling wells with oil-emulsion drilling fluids. These were conventional aqueous muds with a dispersed oil phase. Since 1950, oil-emulsion muds have been used on an increasing number of wells each year. However, the practice of measuring only the resistivity of the mud at the well site has continued, and the mud filtrate and mud cake resistivities have been determined by the above-mentioned relationships. Service companies are now equipped to measure all three resistivities at the well site. An investigation was conducted on the resistivities of oil-emulsion muds, mud filtrates, and mud cakes to determine if these values conformed to the relationships for aqueous muds. TYPES OF MUDS Fifty-one oil-emulsion mud samples were prepared in the laboratory following a standard manual' published by a leading mud company. The diesel oil in the samples varied from 5 to 50 per cent, the majority of the samples being in the 10 per cent region. The basic aqueous mud types which were converted to oil-emulsion muds were commercial clay and bentonite muds, low pH and high pH, caustic-quebracho treated muds, and lime treated muds. The emulsions were stabilized by dispersed solids, lignins, lignosulfo-nates, sodium carboxymethyl cellulose, or sulfonated petrolatum. It is worthy of note that after a quiescent period of two weeks at room temperature all samples, regardless of emulsifying agent, remained stable. The make-up water for the muds was from the laboratory tap. Resistivities were varied by the addition of table salt to the water. A range of mud resistivities from 0.44 to 3.9 ohm-m was obtained in this way. Twenty-three field muds were tested. These covered the same range of mud types as did laboratory muds. Oil provinces of the Gulf Coast, South Texas, West Texas, Oklahoma, Montana, and Canada were represented. MUD TEST PROCEDURE Each mud was tested for density, viscosity, pH, and filter loss by standard testing techniques. The resistivity measurements were obtained with a Schlumberger EMT meter. This meter required small volumes of sample, e.g., 2 mm. Filtrate was obtained from a Standard Baroid fil-ter press at the end of a 30-minute test. The filter cake from the same test was used for cake resistivity measurements. Mud, filtrate, and cake samples were heated to 100" F in a constant temperature water bath prior to measurement of resistivities. RESULTS The relation between mud resistivity (Rm) and mud filtrate resistivity (Rmf) is shown in Fig. 1. The solid line represents an average for the data. The equation of this line is Rmf =0.876 (Rm) 1.075 . . (1) Arbitrary limits, indicated by the dashed curves, have been set. The majority of the data falls within these limits, but some points do lie outside the limits. The approximate equation Rmt = 0.88 Rm , . . . . (2) will give satisfactory results within these limits. The data on mud cake resistivity Rmc is shown in Fig. 2. The solid line is an average for the data. The equation for the line is Rmc = 1.306 (Rm)0.88 The dashed lines are arbitrary limits on the data. Within these limits, Eq. 3 may be simplified to Rmc = 1.31 Rm . . . . (4) DISCUSSION The limiting curves in Figs. 1 and 2 represent maximum deviations of ±25 per cent. Thus the use of the average curves can introduce considerable error. There is no substitute for accurate measurements of mud, mud cake, and mud filtrate resistivities at the well site. The mud sample tested should be representative of the mud opposite the formation being logged. The average mud filtrate resistivity curve of Fig. 1 is reproduced in Fig. 3 with two curves which have been published for clay-base aqueous muds2,3. The latter curves were determined from average values of a large number of drilling fluids. The three curves have essentially the same slope and the differences between them are from 7 to 22 per cent. Comparison is made only to illustrate the possibility of error

Jan 1, 1958

By R. L. Pienaar

Paper written on project work carried out in partial fulfilment of B Eng (Min. Eng.) degree A study was conducted to investigated and compare the feasibility of two single level extraction layouts for an intact part of theorebody of the De Beers diamond mine, Koffiefontein. Two layouts were chosen and the tonnage profile, development schedule and production ring blast layouts were calculated by applying the information acquired through a literature study. Current principles used on the mine were also applied. The first layout chosen is described as a simple straightforward layout with 6 production tunnels spaced 15 m apart, and is referred to as Layout A. The second layout, named Layout B, is described as a complex chevron layout with 9 production tunnels spaced 15 m apart. This layout has a rim tunnel situated in the kimberlite ground, which is called a chevron and this made the layout more complex than Layout A. By comparing the financial results of the two layouts, it was found that Layout B is a more feasible option than Layout A. Sensitivity analyses show that both layouts are extremely sensitive to the revenue of the projects. Therefore it is recommended that Layout B can prove to be a source of additional tonnage to add to the total tonnage profile of the mine and to extend the life of mine by 20 months, only if certain financial/economic factors improve. There is, however, still some areas of the project that need to be investigated.

Jan 1, 2006