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Ninth International Symposium TECHNIQUES FOR DECISION MAKING IN THE MINERAL INDUSTRY The Symposium which is to be held in Montreal during the week of 14th to 19th June, 1970, follows the Computer Symposium held in Salt Lake City in September, 1969. The purpose of the Symposium is to define the role of decision-making techniques within the mineral industry and to stimulate the development of specific applications. The subject of the Symposium-mathematics, operations research and computers as applied to the mineral industry-have permitted us to structure it in such a way that a complete picture of the mineral industry in its public policy, corporate, financial and technological environment emerges. Topics for the sessions are: . The Public Policy Environment;. The role of Decision Making Techniques in the Mining Company; . Information Storage and Retrieval; . Markets-Forecasting, Patterns and Strategies;. Search for Mineral Deposits;. Quantification of Geological Variables;. The Mine Development Decision; Production Systems;. Planning, Scheduling and Control Systems. The sessions will be directed energetically by chairmen who have been in continuous contact with the authors, providing them with information about related papers and sessions. This symposium is sponsored by the Canadian Institute of Mining and Metallurgy, McGill University and Ecole Polytechnique. Further information may be obtained from: The 9th International Symposium. W. H. Gauvin, General Chairman, J. Elbrond, Programme Chairman. Department of Mining Engineering, Ecole Poly technique, P.O. Box 501, Snowdon, Montreal 248, Quebec, Canada. ACHEMA 1970 The European meeting of Chemical Engineering and the Achema Congress 1970 is being held at Frankfurt/Maine, from 17th to 24th June, 1970. The programme includes some 110 plenary and specialised lectures, numerous technical films, excursions, and technical discussions in front of the exhibits which form the central feature of the Achema-Exhibition Congress. Further information may be obtained from the honorary representative of Achema in South Africa, Mr F. Schilling, P.O. Box 31274, Braamfontein, Johannesburg. NINTH COMMONWEALTH CONGRESS PROCEEDINGS Publication of the proceedings of the Ninth Commonwealth Mining and Metallurgical Congress, 1969, is being undertaken by the Institution of Mining and Metallurgy, London. There will be four volumes covering papers presented at the Congress together with the discussion and author's replies, and the addresses by Sir Duncan and Sir Ronald Prain. These will be available April-June 1970. Special pre-publication prices apply to orders received before 1st April, 1970. Order forms may be obtained from Secretary S.A.I.M.M., P.O. Box 1183, Johannesburg. 6th World Congress on the Prevention of Occupational Risks This Congress, dealing with industrial accidents and occupational diseases, will be held in Vienna from 10th to 15th May, 1971. Further information may be obtained from: VI Weltkongress für Unfallverhütung, Postfach 745, A-1015 Wien, ÖSTERREICH. International Vacuum Metallurgy Conference 1970 The Vacuum Metallurgy Division of the American Vacuum Society is organizing an international meeting that will be held at the Convention Center in Aanaheim, California from 15th to 19th June, 1970. Enquiries should be directed to the Arrangement's Chairman, L. M. Bianchi, TRW Inc., 23555 Euclid Avenue, Cleveland, Ohio 44117, Tel. No. (216) 383-2111. AIME WORLD LEAD-ZINC SYMPOSIUM This symposium is being held from 21st to 23rd October, 1970, in St. Louis, Missouri, U.S.A., in conjunction with the Fall Meeting of the Society of Mining Engineers, and a Rapid Excavation Conference. In addition to local field trips to smelter and refinery operations there will be a number of extended trips of six to eight days. Information may be obtained from: Mr. John C. Fox, Secretary, Society of Mining Engineers of AI.M.E., 354 East 47th Street, New York, N.Y. 10017, U.S.A. NEW SUBSCRIPTION RATE TO THE JOURNAL As from 1st JANUARY, 1970, the subscription rate for the Journal has been increased from R12.00 to R18.00 per annum, i.e. R1.50 per issue instead of R1.00. The discount to booksellers is 15 per cent. COURSEOF THE THEORYOF THE UNIT OPERATION OF GRINDING The South African Institute of Chemical Engineers is arranging for Prof. L. G. Austen to conduct a course on the above subject in September, 1970. Interested persons should contact Dr G. S. Harrison, S.A. Institution of Chemical Engineers, A.E. & C.I., 15 Anderson Street, Johannesburg. METRICATION The S.A. Bureau of Standards has published a document entitled "Metric Scales and Units for use in the Building and Construction Industry." It has been compiled by the Metrication Committee of the National Building and Construction Advisory Council and is the first of the series of informatory documents relating to the programme for the metrication of the Building Act Construction Industry. This document is obtainable from the South African Bureau of Standards, Private Bag 191, Pretoria at 50c per copy.

The AK6 kimberlite is situated 25 km south of the Debswana Orapa Mine in Botswana and was discovered by De Beers geologists in 1969 during the follow-up of geophysical targets in the Orapa area. The kimberlite was not extensively pursued at the time as the initial bulk sampling indicated it to be of limited size and low grade, factors largely contributed to by the basalt breccia capping. Completion of high resolution integrated geophysical techniques and drill bulk sampling to depth recovered 97 tons of kimberlite during 2003 and 2004, which led to the increased size and grade estimates. Bulk sampling by Large Diameter Drilling (LDD, 23 inch diameter) commenced in 2005; 13 holes were drilled to a cumulative depth of 3,699 m and 689 carats of diamonds were recovered. In July 2006 the De Beers Mineral Resource Classification Committee classified these Phase I LOO results at a High Inferred level with an average grade of 24 carats per hundred tonnes (cpht) at a bottom cut-off of +1 mm, and a modeled average diamond value of 150 dollars per carat. A second phase of LDO drilling was initiated in 2006, and bulk sampling by trenching commenced in 2007 in order to deliver a resource estimate at indicated level. An Indicated Resource of 11.1 million carats at an average grade of 22 cpht was declared for the deposit mining lease application lodged in 2007.

Jan 1, 2010

By P. F. Sorensen

A simple mathematical model is presented that relates the 'head' developed at the surface of pulp in a tank to the power dissipated by the air in producing it. The value for the head is then used in the calculation of the superficial circulating velocity of the pulp and the area of the overflow, and the product of these two values gives the pumping rate of the pulp. Thus, the superficial circulating velocity, and hence pumping rate of the pulp, are related directly to the flow-rate of the air. This hypothesis was tested with good results in an air-agitated pachuca tank (of 1 m in diameter) fitted with a draught tube. The method can also be used for the prediction of the pumping rates in production-scale vessels.

Jan 1, 1985

By B. G. Meyburgh

SYNOPSIS As a result of a test programme over a period of approximately ten years and a decision to extend the uranium plant, it was decided to design and erect the first full-scale Purlex Plant in the Republic of South Africa. The plant at Buffelsfontein G.M. Co. Ltd. was commissioned in August, 1968, and to date has produced results superior to those predicted in the original estimates. The capital expenditure was R678 000 for a plant capable of treating 1 600 gallons per minute of pregnant solution. Compared with the conventional resin ion exchange process, it has been found that the saving in production costs is 21 cents per pound U3O8 produced. The amine loss is 10 parts per million, which is equivalent to a total organic loss of 250 parts per million. The uranium loss in the raffinate is less than 0.001 g/l U3O8. The final product is of a high quality and it is probable that it is pure enough to convert to the fluoride directly.

By D. Moradkhani, S. M. Javad Koleini, M. Abdollahy, F. Soltani

"The objective of the study was to determine the type and distribution of rare earth elements (REEs) in the phosphate flotation concentrate from the Esfordi flotation plant in central Iran, in order to understand their leaching behaviour. Samples of the concentrate were prepared and analysed using XRF, XRD, ICP, and EMPA. The Ce, Nd, La, and Y contents were 5608, 2227, 1959, and 679.7 mg/kg, respectively. The main REEbearing minerals in the concentrate are fluorapatite, monazite, and xenotime. Apatite is the chief mineral of the concentrate and the ratio of apatite to monazite is greater than 60. From EMPA analysis, monazite contains a much higher proportion of REEs than apatite, but more than 50% of the REE content of the concentrate is in apatite, as this is the main mineral constituent of the concentrate. During the leaching of REEs in the phosphoric acid production process, more than 99% of the total rare earth values report to the large amounts of phosphogypsum (CaSO4·2H2O) formed during the dissolution reaction. Accordingly, two process options are suggested for treating the Esfordi concentrate. In the first option, the concentrate is treated with 98% H2SO4 at 200–300°C. After digestion and water leaching, phosphorus and REEs are precipitated out of solution. In the second option, the concentrate is treated in a pre-leach stage. The process residue from the pre-leach stage is then subjected to acid digestion at elevated temperature for phosphoric acid production.IntroductionThe rare earth elements (REEs) comprise a set of 17 chemical elements in the periodic table, specifically the 15 lanthanides plus scandium and yttrium. REEs and alloys that contain them are used in devices such as computer memory, rechargeable batteries, cell phones, catalytic converters, magnets, fluorescent lighting, and many more (Krishnamurthy and Gupta, 2004). China supplies about 94% of the REE demand, with the remaining 6% coming from Russia and Estonia, the USA, India, Malaysia, and Brazil (Zhanheng, 2011). Increased industrial development in China has prompted the Chinese government to limit annual export quotas to approximately 35 kt of rare earth oxides (REOs), while non-Chinese annual demand is expected to reach 80 kt by the year 2015. This constriction of supply is being met by the development of many new rare earth mining projects, each of which has its own unique mining and processing challenges (Jordens, Cheng, and Waters, 2013). One of these projects is the Esfordi phosphate project, which is located in Yazd Province in central Iran. The concentrate from the Esfordi flotation plant contains 1.2–1.5% REEs."

Jan 1, 2017

By C. T. Connor, O&apos

The fifth University of Cape Town (UC1)/University of Stelenbosch (US) Minerals Processing Symposium was held at the Faculty of Engineering, University of Stelenbosch on Friday, 8th August, 1986. This series was initiated in 1982, the symposia being held alternately in Cape Town and Stelenbosch. The South African Institute of Mining and Metallurgy has kindly agreed to sponsor this event for the next three years. Cokemaking On the initiative of the European Cokemaking Commitee, the First International Cokemaking Congress will be held in Essen (F.R. Germany) from 13th to 19th September, 1987. It is being organized by the Steinkohlenbergbauverein, together with the Verein Deutscher Eisenhitenleute.

Jan 1, 1986

By S. R. Xie, S. Gong, E. P. Li, S. S. He, Y. J. Sun, S. J. Li

"Based on the complicated geological production conditions and abnormal strata behavior of the fully mechanized caving face No. 102 in the Xinyang Coal Mine, we analyze the breaking and forming process of distinguishing key block structures in close spaced coal seams. This analysis establishes an integrated mechanical model of these distinguishing structures. Furthermore, the horizontal force TAB generated by downward key blocks A and B is quantified, as is the Pressure PG exerted by key block B on caved gangue. Finally, the interaction mechanism is revealed. Analysis of the results using Mathematica reveal the following: 1. Both TAB and PG exhibit approximately inverse relationships to coal pillar width b; 2. TAB exhibits an approximately exponential relationship with mining height M of which the base is greater than 1; 3. PG exhibits an approximately arctangent relationship with mining height M. Combining these results with field observations, we conclude that the area in the vicinity of supports 100#–120# in the upper area of working face No. 102 is the likely zone of abnormal strata behavior. Thus, this area was classified as a key control area. Based on these results, the graded support control technique is proposed. Application of this technique in the field resulted in a reduction in the rate of hydraulic support system malfunction from 47% to 7% in the key control area. In addition, during times of periodic mining pressure, the roof caving height value never exceeded 0.3 m and the maximum spalling depth never exceeded 0.2 m. Overall, the results indicate that this new technique has markedly increased the stability of surrounding support rock material. IntroductionClose-spaced coal seams are widely distributed in China, and mining proceeds from the uppermost to the lowermost seam (Zhang et al., 2008). In the case of coal seams in close proximity, a particular set of risk factors becomes pertinent. These include disturbance of lower seams during the mining of overlying ones and the decrease in mechanical stability of lower coal seams following repeated movement of key blocks in main coal seam roofs. This inevitably results in behaviour such as heavy wall spalling and support damage to fully mechanized caving faces in lower-lying coal seams, which necessitates a significantly increase in the measures to support surrounding rock material (Zhang et al., 2010; Huang et al., 2013; Xie et al., 2013)."

Jan 1, 2016

One of the main bottlenecks in the carbon-in pulp process for the recovery qf gold is the elution procedure, which typically requires the heading of caustic cyanide eluants to high temperatures for 16 to 24 hours. The present work demonstrates that sodium sulphide solution as an alternative eluant can effect complete elution in about 4 hours at ambient temperatures. Elution efficiencies of around 100 per cent were obtained in 4 hours with a single pass of eluant containing 0,2 M Na2S and 0,4 NaOH about 10 bed-volumes of eluant The initial rate was slow over the first hour of elution, probably because the activated carbon catalysed the oxidation of sulphide to polysulphide. Elution efficiencies of around 100 per cent were also obtained in less than 4 hours during the batch elution of carbon at liquid-to-solid ratio of about 100. Lower liquid to solid ratio resulted in the re-adsorption of gold, probably owing to the oxidation of sulphide to polysulphide, with the resultant formation of gold complexes that were eluted less readily. Improved rates of elution were obtained at higher sulphide concentrations and pH values greater than about 13. Higher temperatures increased the initial elution rate, but lowered the overall extraction efficiency, probably because of the deposition of elemental gold on the carbon. Variation of the ionic strength by the addition of NaCl had no defect on the elution, which confirms that the elution mechanism in the case of sulphide is different from that when cyanide or I1)1droxide is used as the eluant. It is proposed that the elution of gold by sulphide solutions proceeds by means of an initial step that involves the reaction of polysulphide ions with the adsorbed aurocyanide speeds, forming AuCN on the carbon and thiocyanate in solution. This step is followed by the formation of poorly adsorbed complexes with sulphide ions, such as [AUS3-2]-The presence of polysulphides, whether generated in situ by the catalytic oxidation effect of activated carbon or by the addition of elemental sulphur, reduces the elution rate and efficiency dramatically. This is probably due to the formation of complexes such as AUS-3 and AUS-2 , which have a high adsorption affinity.

Jan 1, 1994

By Oaitse Medupe

The feed to the BCL concentrator is made of the ore from Phikwe No3 shaft, Selebi shaft and Selebi North shaft. The receiving bin of the concentrator has daily throughput of 7869mt live capacity of ore with feed grade from underground operation of 0.70%Ni, 0.75% Cu and 0.05% Co. The valuable minerals in the ore are Pentlandite, Chalcopyrite. Cobalt is present mainly in the solid solution in pentlandite. The common non-sulphides minerals are Amphibolites, quartz, micas, and Feldspar while pyrite and pyrrhotite are the major gangue sulphides. Pyrrhotite contains between 0.1% and 1% Ni in solid solution. From the mineralogical investigation, it was found that approximately 65% of the pentlandite in the ore occurs as exsolutions in the pyrrhotite and the grain sizes ranges from 8 to 38µm. about 25% of the nickel bearing pentlandite is medium grained and subhedral. This kind of pentlandite occurs along the cracks of pyrrhotite or as inclusions in pyrrhotite. The rest of pentlandite present is disseminated in the gangue matrix. Because of the fine occurrence of pentlandite, fine grinding was conducted in order to improve the recovery as well as the concentrate quality. A series of subsequent flotation test were performed in order to evaluate grinds of 35, 40, 45, 50, 60 and 80% passing 75µm. It was observed that ? The mass pull increased together with increasing fineness of grind ? Similar nickel recoveries were obtained at 35 and 40% passing 75µm but from 50% and 60% the grind was seen to adversely affect the nickel recovery. At 80% passing 75µm the recovery increased by small amount. ? Flotation kinetics showed higher initial nickel recoveries but these did not translate into higher final recoveries. ? Grinding finer than 80% passing 75µm yield very high grades of the concentrate in the first minute of the flotation test, and the grade is dramatically reduced in the fifth minute of the test. It was concluded that the current grind of approximately 48% passing 75µm is close to the optimum and no significant benefit were found to be associated with finer primary grinds. The inconsistence of the grade and recovery as well as the pentlandite losses will be discussed. Keywords: Pentlandite, Nickel losses, Grain size, Flotation

Jan 1, 2007

By L. A. Vermeulen, C. P. Campbell, D. D. Howat

The production 01material liner than 75 µm in a number of open-circuit mills was investigated. The mils ranged from laboratory rod and ball mils treating a variety 01 ores to large industrial rod mils treating gold and copper ores. It was shown that the dependence on feed rate 01 the production 01 material liner than 751J.1T1 can be accurately described by a simple formula that was derived in previous work. In some 01 the systems reviewed, it proved possible to assess the feed rate 01 ore to the mil at which the rate 01 production 01 material finer than 75 µm can be maximized.

Jan 1, 1991

1. Book review • World mineral statistics 1974 to 1978. Production: Exports: Imports. London, H.M.S.O., Institute of Geological Sciences, 1980. (Reviewer: R. P. King) This edition is the third in a regular series prepared by the Institute of Geological Sciences. It is a useful compilation of statistics on the production, export, and import of fifty-eight different mineral commodities. Each mineral is listed by country, and a table is provided for the annual production, the annual imports, and the annual exports for each commodity included. However, some commodities are treated in considerably more detail; for example, aluminium and bauxite have tables under the following headings: Production of Bauxite, Production of Alumina, Production of Primary Aluminium, Exports of Bauxite, Alumina and Aluminium, Imports of Bauxite, Alumina and Aluminium. The information presented is the latest official information available from primary sources; the compilers have not relied on data from secondary compilations. This publication will be of considerable assistance to all who are concerned with the trading of mineral commodities and their production.

Jan 1, 1981

By F. P. Agterberg, A. G. Fabbri

"A statistical analysis was done on the locations of 718 copper and 1 257 gold deposits in Abitibi. This area of about 20 000 square miles is one of the better-explored metal-producing areas of Canada.Patterns of clusters were studied by harmonic analysis. Several two-dimensional power spectra showed a number of distinct peaks. Sets of crest lines for patterns of interfering waves were constructed on the basis of these peaks. They pass through elongated clusters of deposits which are spaced more or less equally. For copper these periodic regularities in location may coincide with geological structures largely of Archean age (major fractures and sedimentary belts indicating synclinal structures). A directional relationship with swarms of diabase dykes of known ages and distinct attitudes indicates that the gold mineralization is, in part, younger than the copper mineralization."

Jan 1, 2014

By Y. C. Kim

The estimation of local recoverable reserves using the widely spaced exploration drill hole information in an open-pit mine is probably one of the most difficult tasks in ore reserve estimation. The difficulty of this problem is magnified if the grade distribution is highly skewed, such as in a gold deposit. This paper describes the results of performance comparison of three different kriging techniques in an open-pit gold - silver deposit. The three different kriging techniques compared in their local recoverable reserve estimation capabilities are: (1) lognormal kriging, (2) indicator kriging, and (3) probability kriging. For performance comparison, the predictions are compared with the actual recoverable reserves determined from blast hole information. Based on the results from one test bench, both indicator and probability kriging performed better than lognormal kriging, perhaps owing to samples not being truly lognormal. Between indicator and probability kriging, probability kriging results were definitely superior to indicator kriging results.

Jan 1, 1987

By R. T. White

The usual figures quoted for water usage in South African gold mines are 1 ton of water used per ton of rock broken, which represents a consumption of 4000 litres of water every second. Owing to the limited sources under ground water, the high cost of surface-water supplies, and the environmental problems of effluent disposal, water IS recycled continuously for underground use. This paper reviews the various types of water treatment and why water needs to be treated. The uses of lime and some flocculants are considered, since much of the mine service water requires only simple neutralization and clarification. Where mining depths necessitate the installation of underground refrigeration plant, more stringent water requirements must be met, and where plant performance is to be maintained, the prevention of system scaling and corrosion is necessary. The families of chemicals added to provide such protection are outlined, together with a description of the mechanisms of control.

Jan 1, 1985

By E. J. Magri

For life-of-mine planning, as well as for short- and medium-term planning of grades and mine layouts, it is extremely important to have a clear understanding of the patterns followed by the distribution of gold and uranium within the mining area. This study is an attempt to reconcile the geostatistical approach to the determination of ore-shoot directions, via an analysis of the spatial distribution of gold and uranium values, with the sedimentological approach, which is based on the direct measurement of geological features. For the routine geostatistical estimation of ore reserves, the Hartebeestfontein gold mine was divided into 1I sections. In each of these sections, the ore-shoot directions were calculated for gold and uranium from the anisotropies disclosed by geostatistical variogram analyses. This study presents a comparison of these results with those obtained from direct geological measurements of paleo-current directions. The results suggest that geological and geostatistical studies could be of significant mutual benefit.

Jan 1, 1978

By M. I. Mabala, B. P. Watson, M. J. Kanda

A failed and stable Upper Group 2 chromitite pillar database was collected in 2007 as part of the PlatMine 1.2 pillar strength project. The database consisted of 177 pillars, of which 33 had failed. Most of the data were collected from various intermediate-depth mines near the town of Thabazimbi. Spans between rows of chain pillars were about 30 m. Pillar loads were estimated using pseudo-3D, elastic, boundary element modelling software. The pillar-load data were evaluated using a maximum likelihood regression analysis to determine strength parameters, following a similar approach to historically acceptable practice. The original formula was named the PlatMine Formula. This paper describes re-evaluations of the database, using a linear formula and several variations of the power formula. The formula providing the lowest standard deviation to the data in the database was considered the best representation of the pillars in the database. Through this rigorous process, a new formula for peak pillar strength was developed for Upper Group 2 chromitite pillars, which is unaffected by volume. The small standard deviation provided by this formula suggests a reliable relationship between strength and w/h ratio, in the w/h ratio range between 1.5 and 4.7. Accuracy, precision, and recall analysis confirmed the efficacy of this equation to the data in the database. Importantly, the new formula provided in the paper does not predict a significantly different strength to the current PlatMine formula, within the confines of the database.

Nov 27, 2024

[The following notations will be used: O an ore body. z a point in O. z' another point in O. w0 (z) a point support of infinitely small dimensions w0 centred in z. w (z) a sample (or block) of finite dimensions kv centred in z. W (z) a block of size W greater than or equal to w. W'(z) a block of size W' greater than or equal to W. W" (z) a block of size W" greater than or equal to W'. x (z) value of the regionalized variable x with support wO at point z. p (z) average value of x in w (z). pw WZ) = E[x(z')l zt in w(z) pw (z) average value of x in W (z). pwl (z) average value of x in W' (z). pwU (z) average value of x in W" (z). Note that pW (z), pw (z), pwf (z) and pw,, (z) are (regular- ized) regionalized variables with support w, W, W' and W" respectively. The relative sizes of wO, w, W', W" and O are represented schematically in Fig. 7.1. The variance of pw (z) when w (z) takes all possible positions in O is known as the dispersion variance of w in O and is denoted by a? (w in O) or a\: 0; = a2(w in a = Ew,z, inn {[p~ (z) - p121. (7.1) For convenience, the following simpler notation will be used: a> Ew in n [(p w - pl21. Consider now a large block W (z) in O with average value p~(z), and all the possible smaller blocks (or samples) w (z') of size w in W (z). We can calculate the variance of the blocks w (z') in the block W(z), which is the dispersion vaciance of w in W (z), from the relationship a2 [w in W(z)l = Ewcz,, in W(Z) {[PW (z') -pw (z)l21. (7.2) Provided the intrinsic hypothesis is satisfied (§3.4.4), this variance is a function only of the dimensions of the supports w and W, and is independent of the position z of the block W (z). We can therefore define the dispersion variance of w in W, as follows: u2 (win W) = Ew in w [(ptti - PW)~]. (7.3) In cases where the intrinsic hypothesis is not satisfied, the variance of samples w in blocks W may vary only slightly from block to block, so that the average dispersion variance of w in W (z) for all possible W (z) in O can be used: ayw in W) = Ew(,, ~n n {a2 [W in W (z)]}. (7.4) The following example will help to illustrate the importance of the variance of dispersion. Assume that we plan to develop a bedded coal deposit Q. We consider mining the ore body with only one giant dragline, which will extract each day a block of coal of size W'. The total area mined in a year will have a size W". We want to know whether this scheme is acceptable, given that the daily variation in ash content must not exceed a given percentage specified in the sales contract, which is renewable on a yearly basis. On a given day, if we mine the block W' (z), the ash content will average pwT (z). In a period of one year, the expected daily variation in ash content is measured by the expected variance of pwf (z) in W", i.e. by a2 (W'in W"). An alternative mining method might be to use two smaller draglines, each one mining W'/2 of ore per day, and Wn/2 per year. If the draglines are used to mine two distinct sections of the deposit, the expected daily variation in ash content will be u2 (W'/2 in Wj1/2)/2 Other situations, where economic decisions cannot be made without knowledge of the dispersion variance, are illustrated in § 7.4 and 7.5 below. Remark: The notation a2 (w in W) is used in this publication because of its clarity. An alternative notation, which is increasingly accepted in the literature, is s2w/W. ]

Jan 1, 1978

Measurements of the rates of penetration at various throttle air pressures are presented as a function of thrust for three different rockdrills when drilling quartzites with strengths of 45 000 lb/in2 and 39 000 lb/in2, The rate of penetration at optimum thrust is shown to agree well with the predictions of theory and curves of air consumption as a function of throttle air pressure are presented. Man-held machines are shown to be under-thrusted at throttle air pressures above 40 lb/in2g. Measurements of bit wear as a function of distance drilled are presented graphically and discussed in terms of the theory developed in Part 1.

By C. P. S. Barnard, D. G. Krige

Objective justification on economic grounds for expenditure aimed at improving the quality of the underground working environment requires a suitable measure of any improvement in productivity that would result under routine working conditions. For this purpose, a composite production criterion is suggested, as well as its correlation under actual operating conditions with various factors directly affecting productivity and the environment. Of the environmental factors, wet-bulb temperature appears to be the major factor affecting productivity. The production data given indicate that a decrease in the wet-bulb temperature of 1OC can be expected to result, on average, in some 3 to 4 per cent increase in productivity; in the higher temperature ranges, this percentage could be much higher.

Jan 1, 1981

By G. R. Addie

The Hydraulic Institute has completed the task of developing a new ANSI/HI standard 12.1-12.6 (2005) for rotodynamic (centrifugal) slurry pumps covering nomenclature, definitions, applications, and operation. The standard provides examples of the different slurry pump types and contains an extensive section on pump and slurry definitions. The effect of slurry on pump performance is covered along with the pumping of froth. Reference is also made to ANSI/HI standard 9.6.7 (2004), which contains a new method for pump performance correction when handling viscous fluids. Classification of slurry services is established and then is used to determine limitations on velocities and total head per pump in order to obtain acceptable wear performance. The new service class, head per stage and other limits are directly related to capital and other cost considerations that will affect solids transport system economics. The writers review the contents of the new standard, highlight the main points, and discuss the reason for the slurry classification, corresponding limits and expected implications, particularly with respect to operating costs of the pumps in solids transport systems.

Jan 1, 2007