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By Rossen Halatchev

The production sequence optimization of open pit mine is a procedure that reflects the exploitation conditions over life of mine. These conditions are determined by the geological characteristics of the deposit, requirements of the mining and processing technologies and economic parameters of the exploitation. From a technological point of view the production sequence optimization deals with two aspects - spatial and time. Each of them plays a tremendously important role in searching for the optimum solution of the problem. This paper treats the time aspect of the production sequence optimization. This aspect is revealed in the context of the problem of the working zone formation of the open pit mine in time. The problem formulation places an accent on the determination of the theoretically possible time distribution of the waste and ore quantities of each bench of the open pit mine. A mathematical solution to the problem has been developed based on combinatorial analysis. The solution provides a multi-variant investigation of the production sequencing with the use of an optimization model developed as a linear programming formulation. The model objective function uses the Net Present Value as a criterion for assessing the effectiveness of production sequencing. A case study is provided for the conditions of Assarel copper open pit mine (Bulgaria) to illustrate the applicability of the approach. The results obtained reveal the complex nature of the variation of the Net Present Value as a function of the mining factor.

Jan 1, 2002

By C. L. Klingman

In extracting helium from natural gas at Bureau of Mines helium plants, a continuous analysis of the helium content of inlet and outlet gas streams is required to determine extraction efficiency. An analyzer, using established principles of gas chromatography to effect the analysis, has been developed by the Bureau of Mines to meet this need. It will record the helium content of inlet or discharge gases from helium-extraction plants every five minutes, continuously, without the attendance of an operator. A 25-inch column of activated charcoal is used as a partitioning agent; nitrogen is the carrier gas; and a hot-wire, thermal-conductivity cell is the detector. A 1.0 mv, strip chart indicates the unbalance of the detector cell, Sample size is 0.37 cc, and carrier gas flow rate is 75 cc, per minute. Significant improvements in helium extraction have been demonstrated by using the chromatograph in three helium-extraction plants.

Jan 1, 1961

It is possible to closely simulate the product sizing distribution from an industrial rod mill by the correct choice of parameters in laboratory batch milling. For laboratory batch rod milling, changes in per cent critical speed and per cent rod volume cause the product sizing distributions to move in parallel at a constant slope. For ball milling, larger sized media in batch milling grind more coarse material giving a closer simulation to the industrial rod mill product. Changing the feed size distribution by removal of coarse particles or addition of fine particles changes the product sizing distribution, but would not normally be practiced in laboratory testing. These data should be applicable to many instances where laboratory simulation of an industrial product size distribution is required.

Jan 1, 1998

By John H. Soule

"Although known since 1902, the Capitan iron deposits have not been mined, mainly because of the low grade of the ore. Only a small amount of exploratory work has been done since the discovery. C. H. Johnson, district engineer of the Bureau of Mines in New Mexico in 1942-45, visited the property; as a result, preliminary sampling of the property was done in 1943.3/ Owing to the results of this preliminary work, an extensive development program by wagon drilling was planned and executed early in 1944.4/ Equipment available at that time was limited in the depths that could be reached; and, as geologic studies indicated the possibility that other limestone beds below the surface might have been replaced by iron minerals, a supplementary development program by churn drilling to reach greater depths was prosecuted in late 1947 and early 1948, with the author as project engineer.The following sections note the location and accessibility of the deposits and mention the physical features, climate, labor and living conditions. Property and ownership are outlined, the history and production of the property stated briefly, and the ore deposits are described. The work done by the Bureau of Mines is explained and supported with maps, sections, and assay logs.From March 1, 1944, to June 1, 1944, 179 wagon-drill holes totaling 3,490 feet and 4 test pits totaling 75 feet were completed, and from November 1947 to February 1948 seven churn-drill holes totaling 2,488 feet were drilled and sampled. Metallurgical tests are summarized and referenced.With respect to this report, special acknowledgment is due to J. H. Hedges, chief, Tuscon, Ariz., Branch, Mining Division, and to W. R. Storms, engineer in-charge of the Silver City, N. Mex., field office. Acknowledgment also is due to J. Bruce Clemmer, chief, Tuscon Branch, Metallurgical Division, for analyses done in the last phase of this project, and to Dr. Vincent Kelley of the Federal Geological Survey for his cooperation during this work.The Capitan iron deposits are in secs. 10, 11, 13, 14, and 15, T. 8 S., R. 14 E., central Lincoln County, N. Mex., as shown on figures 1 and 2. The only known large deposit upon which the Bureau did development work lies around the common corner of secs. 10, 11, 14, and 15, as shown in figures 2 and 3, and just southwest of and in the foothills of the Capitan Mountains. This range of mountains trends east and west. The elevation of the highest peak is 10,205 feet.The deposits are reached by traveling on State Highway 48 from Capitan north for 4 miles, to a point just inside the Lincoln National Forest fence and thence due east over a poor dirt road for 1.2 miles. Just over a small wooden bridge, the left-fork is taken, and the poor road is followed to the deposits, a further distance of 1.1 miles (see fig. 1). This road follows the fence to and through the Davis ranch. All the roads from Capitan to the deposits are dirt roads but are impassable only for short periods during and after heavy rain or snow."

Aug 1, 1949

By Ralph Sherman

IN these days of intense competition for the domestic fuel market much is heard of the low efficiencies with which solid fuels, and most particularly bituminous coal, are burned in domestic heating equipment. In contrast to the every-day operating efficiencies of 85 and 90 per cent in large, central, electric-generating stations, efficiencies of 50 per cent and lower are cited for the hand-fired home furnace. Just as the miles per gallon of gasoline is only one of the factors that determine the cost of operation of an automobile, so the thermal efficiency is only one factor in the cost of domestic or other heating. The purpose of this paper is to analyze the total costs of domestic heating by various fuels and types of equipment, to determine the relative importance of efficiency and other factors, and to discuss what this signifies relative to the future of the use of fuels in the home.

Jan 1, 1933

By H. M. Faust

THERE are a number of different types of power plants. Some use coal to generate steam, others use oil or gas either in the form of steam or internal combustion equipment, while still others avail themselves of water power as a source of energy. A general comparison of the relative power production with different sources of energy in the United States, for several different years is set up in Table 1. This table shows that coal was used to generate 85.6 per cent of the total heat power requirements of 1909, and only 59.2 per cent of the 1930 requirements. The increasing proportion of modern heat energy generated by the use of fuel other than coal is apparent. If coal is to maintain its importance as a source of heat energy, the necessity for its improved perform-ance, in which preparation plays an -important part, cannot be stressed too strongly.

Jan 1, 1933

By E. Keskinkilic

In this study, sulfide-type inclusion morphologies of a Ca-treated hot-rolled wheel steel, Grade-X, were investigated. The specimens taken from the hot-rolled products of Grade-X heats were studied. The size and distribution of elongated sulfide inclusions were examined for the heats of different S levels. Elongated sulfide inclusions were either not observed or rarely encountered when ladle analysis S content was below 50 ppm. The size and relative abundance of sulfide inclusions were reported to increase markedly when the S composition was greater than or equal to 70 ppm and the Ca/S ratio was less than 0.50. The effect of Ti addition (0.010?0.020% Ti) on sulfide-type inclusion morphologies was then investigated for heats having a sulfur content higher than 50 ppm. Addition of Ti was found to have a partial effect on decreasing the size of sulfide-type inclusions under real steelmaking conditions. Keywords Sulfide inclusions, Ca-treatment, titanium.

Jan 1, 2011

By Cudmore JF, Rowlands D

Incidents associated with spontaneous combustion of coal have resulted in the loss of many lives and considerable capital and equip- ment. Nevertheless, the factors controlling the propensity of coal to spontaneous combustion are not clearly understood and there is a need for research to study the effects of rank, petrog- raphy, moisture contents, etc. This has been attempted in a study of the propensities of selected Queensland coals to spontaneous combustion. An adiabatic calori- meter was constructed to study the self-heating of a large number of coals. Chemical and petro- graphic analyses of the coals were determined and an attempt was made to relate these propert- ies to the self-heating characteristics of the coal. The results show a strong relationship between rank and self-heating characteristics and that petrographic composition of coal may also be of importance. Some observations on the application of the laboratory results to the practical situation are made as are recommendations for future research.

Jan 1, 1981

By D. M. Walters

To achieve safe and healthy working conditions in shaft sinking operations, ventilation, and where necessary cooling, must be provided. These requirements are considered for sinking vertical, sub-vertical and decline shafts. Regulatory requirements, determining air volumes and amount of cooling required, the various ventilation methods used such as force, exhaust, twin columns, air reversal using four gate systems and the equipment required to achieve this, are discussed. Consideration is also given to the sinking of single and twin declines using diesel equipment, calculating the air volumes required to dilute pollutants and the various methods for achieving this. Also discussed is the importance of having early ventilation design input to the overall design of the mine. In many cases ventilation can be the prime determinant for the size and configuration of the shafts or declines that have to be sunk.

Jan 1, 2009

By D. B. Robinson, J. Lukacs

A knowledge of the equilibrium water content of hydrocarbon systems under pressure is important to the natural gas industry. The information available on the solubility of water in hydrocarbon, hydrogen sulfide, and carbon dioxide systems is reviewed in this paper and the influence of the more important variables such as temperature, pressure and molecular structure on solubility in liquids and gases is discussed. A suitable cbromatographic technique has been developed for determining low concentrations of water. Tailing of the water peaks has been elim inated by adding water to the carrier gas stream The method is applicable for both gas or liquid samples and is effective in the presence of hydrogen sulfide. The experimental study of water solubility in methane-hydrogen sulfide systems at a temperature of 16oF has shown that the presence of hydrogen sulfide causes only a modest increase in water content at pressures up to 1,400 psia. Theoretical considerations and data on pure hydrogen sulfide and carbon dioxide suggest that the effect of both these compounds will be greater at higher pressures and in the liquid phase. INTRODUCTION Before transporting or processing natural gases and gas condensates, it is usually necessary to dry them using suitable dehydration equipment. The design and operation of this equipment requires a knowledge of the amount of water present in the fluid at the reservoir and operating conditions. This is influenced by temperature, pressure and composition, particular1y when certain non-hydrocarbon components are present. Field experience indicates that hydrogen sulfide and carbon dioxide, for example, alter the usual water solubility relationships appreciably. However, an extensive search of the literature does not reveal any quantitative data on such systems. For sweet natural gases, generalized empirical correlations such as the one proposed by Katz, et al,6 can be used to predict water solubility with confidence at most temperatures and pressures of interest. However, existing theoretical relationships do not permit a calculation of the deviation from these curves when polar substances like hydrogen sulfide are present in the system. Thus one must resort to an experimental approach to obtain the necessary information. The fact that laboratory experimental methods frequently involve the use of mercury which reacts with hydrogen sulfide in the presence of water, and that hydrogen sulfide interferes with many chemical reactions specific for water has contributed to the difficulty of studying water solubility in systems containing hydrogen sulfide. In this investigation the water content of a limited member of methane-hydrogen sulfide mixtures was determined using a special technique with gas chromatography. REVIEW OF PUBLISHED DATA Experimentally determined water solubility data have been reported for methane, 14 ethane, 15 propane, 9 n-butane, 16 1-butene, 11 hydrogen sulfide 18 and carbon dioxide. 19 These studies report the effect of pressure, temperature and molecular structure on water solubility in single component gases and liquids. Fig. 1 compares experimentally determined water solubilities in several hydrocarbons, carbon dioxide and hydrogen sulfide at l00F as taken from this

By William Petruk

Performance assessments are carried out to monitor operations, to evaluate performance and to identify processing problems. The assessment involves collecting and analyzing representative samples from critical points in a flowsheet and interpreting the data. Sample analysis techniques include chemical analyses, mineralogical investigations, mineralogical analyses, and materials balancing. Chemical analyses provide the lowest cost assessment and should be used liberally, particularly in monitoring a process. Materials balances performed on chemical assays for a suite of samples are used as a first step in evaluating a process. Mineralogical investigations are more expensive than chemical analyses but need to be used only at the beginning of mining and metallurgical operations and as a second step in assessing processing problems. Quantitative mineralogical analyses combined with materials balancing provide the most complete performance assessment, but the procedure is expensive. On the other hand, the assessment has the potential of a large payback when carried out where warranted.

Jan 1, 1991

By A. F. Crawley

The density and viscosity of 1iquid thallium have been measured by absolute methods to temperatures of about 200° and 150°C, respectively, above the melting point. These new data reported, especially density data, do not closely confirm previous work. Density p, in g per cu Cm, is shown to vary linearly with temperaluve t, in °C, according to the equation p = 11.658 - 1.439 X l0-3t. The viscosity data obey the well-known Andrade equation nv1/3 = A exp C/vT , the constants A and C for thallium having values of 2.19 x A and 79.648, respectively. This paper reports some new data for the density and viscosity .of liquid thallium. Measurements of these fundamental physical properties were undertaken as part of a continuing research program at the Mines Branch, Department of Energy, Mines and Resources, Ottawa. Canada. A literature search has revealed that data are so scarce that there could not be a consensus on the true values of the density and viscosity of liquid thallium. To be more specific, there exists only one set of viscosity data' and only two acceptable sets of density data,273 one of which is limited in scope.3 In Liquid Metals Handbook,3 another density study is reported but indications of impurities in the thallium render the results suspect. In this situation, further careful experimentation was required to realize the true density and viscosity of thallium. EXPERIMENTAL METHODS Density. Densities were determined using a graphite pycnometer. The technique and its accuracy have been discussed in earlier papers.4'5 It is considered that experimental data can be obtained which are accurate within +0.05 pct, all sources of random and systematic errors having been evaluated. Density results for thallium were identical whether measured under an atmosphere of argon or a vacuum of 5 x 10-6 torr and, for the most part, the argon atmosphere was used. Viscosity. Viscosity measurements were made in an oscillational viscosimeter by an absolute method—the liquid metal being held in a closed graphite cylinder. Design and operation of the apparatus, constructed in this laboratory, have previously been discussed.6 For thallium, runs were made under a vacuum of about 2 x 10-6 torr. To evaluate viscosity coefficients from the various experimental parameters, the mathematical analysis of Roscoe7 was used. Measurements of the necessary parameters and the accuracy of these measurements have also been discussed.6 The cylinder dimensions were corrected for the anisotropic expansion of graphite, as discussed for density measurements.4,5 It is well-known that thallium oxidizes rapidly and hence a newly machined surface quickly tarnishes in air. The oxide film, however. is nonadherent and is easily removed by rubbing or by solution in water. Hence, immediately before use, both density and viscosity charges were immersed in water, wiped dry, and quickly transferred to the apparatus which was then rapidly evacuated. Specimens removed after determinations were only slightly tarnished and there was no other evidence that tarnishing affected the results. For example, the sharpness of the specimen edges from the containing vessels indicated complete filling by the liquid metal. Thallium of 99.999 pct purity was used in this investigation. Because of its high toxicity care was exercised in handling this material. For example, the melting procedure to prepare machinable ingots was carried out in an open, well-ventilated area, while protective gloves were always worn when handling the solid metal. RESULTS AND DISCUSSION Density. Measurements were made over a tempera-ture range of about 200°C above the melting point. The results are listed in Table I and plotted in Fig. 1. From the graph it is evident that the relation between density and temperature is linear. Such a relation has been observed before in this program for other metals and alloys475 and elsewhere by other workers. A least-squares analysis of experimental data gives the equation: pT1 = 11.658 - 1.439 x 10-3t where p = density in g per cu cm and t = temperature in "C. In Fig. 1, together with the present results, the data of Schneider and Heymer2 in the corresponding temperature range have also been plotted. Evidently, the two sets of data do not agree well, the results of Schneider and Heymer being about 0.6 pct higher. Viscosity. Viscosity data were obtained from the melting point, 303.5°C, up to 457.5"C. The data are listed in Table I and in Fig. 2 the plot of these results demonstrates a smooth curvilinear relation between

Jan 1, 1969

By H. M. Chance

WITHIN the last 10 years the subject of mine taxation in its relation to coal-mining interests has come to have growing importance, not only to those engaged in the mining of coal, but also to the owners of land underlain by coal. A disposition has developed among legislators to accept the principle, which has been laid down y many judicial decisions, that coal underlying the surface must be regarded as a part of the real value of all properties so underlain, and as such is chargeable with such burdens of taxation as the necessities of the government may impose. These necessities arise from the normal expense of maintaining State and local governments, and in the past have been met principally by, taxes based upon assessments of real estate values, which values in many cases have been confined principally or wholly to the value of the surface, without taking into consideration the value of minerals underlying the surface. It is not proposed here to deal with the question of Federal taxation, whether levied on production, or on products transported from one State to another, or exported from the country, or upon thegross or net earnings of corporations engaged in interstate commerce, or upon the production of materials which may be transported from one State to another, or levied. upon the profits or upon the obligations of corporations. The present discussion will be confined to the appraisal of coal lands as assessed for State, county, township, or borough taxation. The methods heretofore and at present in use in the different States vary widely and have resulted in more or less litigation. Discussions appearing from time to time in the technical and trade journals and in the proceedings of mining and engineering associations frequently direct attention to the need for the adoption of fixed principle in levying taxes, and of uniform methods for applying such principles, but there does not seem to be a general demand from the public for the adoption of such methods.

Jan 7, 1914

By Donald Leroux, Dave Scott, Giovanni Di Prisco

"The Lady Loretta lead-zinc deposit in Northwest Queensland, Australia contains an average of 5.9% lead, 17.1% zinc and 97 g/t silver. Ore mineralization is mostly comprised of fairly complex intergrowths of pyrite, sphalerite and galena. Starting in 1998, Noranda Inc., the major joint venture partner of the project, conducted extensive metallurgical test work, which culminated in July - August 2000 in a series of Pilot Plant tests. A common thread to all the metallurgical work has been the deportment of a large portion of silver into non-payable metallurgical products. Average silver distribution from four Pilot Plants shows a distribution of silver in lead concentrate of 28%, whereas silver in zinc concentrate was 49%, and 23% in the final tails.The consistently poor silver recovery as well as the fairly unusual silver distribution, particularly in zinc concentrate, was considered an issue to be thoroughly investigated. A mineralogical plan was devised to determine if these particular metallurgical results were somewhat related to the nature and mode of occurrence of silver carriers. Work was performed to determine discrete silver-bearing minerals and/or silver contained in solid solution in a suite of commonly occurring sulphides. Ore microscopy, SEM methods, Electron Microprobe and Ion Microprobe analyses were all used to answer these questions, and to find the silver carriers, so a metallurgical plan could be devised to improve silver recovery.INTRODUCTIONPotential economic exploitation of the Lady Loretta deposit in Queensland, Australia is driven by revenue derived from zinc and lead flotation concentrates and co-product silver recovery in lead. Relatively low recoveries of payable silver to lead concentrate, and high silver distribution to zinc concentrate necessitated an investigation into the natu"

Jan 1, 2002

By M. J. H. Ruscoe

"The circulation wear rates of pure nickel, cupronickel and nickel-plated-steel coins were measured and found to be in the ratio 1:3:1 . The relative wear rates of nickel and cupronickel coins were similar to those reported in earlier studies, and the circulation wear data for nickel were used as a reference in laboratory wear tests. A correlation between circulation wear rates and laboratory wear rates formed the basis for a prediction of service life using relief legibility as an 'end of life' criterion. This criterion was determined to be an average surface thickness loss of 12 um. The surf aces of nickel and Nickel-Banded-Steel (N-B-STM) coins were observed to harden both in circulation and in the laboratory tests at a rate of approximately I VHN/year of actual or simulated service life. It was also observed that the laboratory wear rate of N-B -S coins increased as nickel cladding thickness increased and this effect was attributed to differences in surface hardness.Wear test results are also reported for the new Canadian dollar coin material which shows a better wear resistance than pure nickel. IntroductionThe subject of coinage wear has a long history, and a number of studies have been published, concerned both with the measurement of circulation wear rates and with accelerated laboratory tests to simulate the behaviour of coinage materials in service. As early as 1803, wear tests were conducted by Hatchett on gold coinage using various types of vibratory and tumbling machines». These tests were commissioned by King George III of England who was concerned about the wear of gold coins in circulation.Early in this century, Adrian performed laboratory tests that showed that the wear rates of cupronickel 75/25 and nickel coins were in the ratio 3:1 and this ratio was later confirmed in laboratory tests reported by Daeppen and Zurrerw, The latter authors also weighed cupronickel and nickel coins that had been in circulation for various times and observed the same 3: I ratio. There was further support from Tylczak and Blickensderfer who reported significantly higher abrasive wear rates for cupronickel compared to nickel, when rubbed against sand paper."

Jan 1, 1988

"Acta Metallurgical, C E Birchenall, Dec '58 Alfred Noble Joint Prize Committee F B Foley, Dec '61 American Association for the Advancement of Science J T MacKenzie, Dec '58, 0 B J Fraser, Dec '57 American Geological Institute Charles F Park, Oct '58, Arthur Montgomery, Oct '59 Committee on Government Relations American Power Conference J P Weir, T D Hess American Standards Association , Mining Standards Board E R Borcherdt, Jan '59, Otto Herres, '59, J K Richardson, Dec '58, Alternate E W Bullard

Jan 1, 1958

By Jeff W. Butwell

Mill flotation tailings impounded at the Gooseberry Mine site between 1976 and 1984 were determined to be an economic source of gold and silver. The tailings would be recoverable by standard heap leach technology in conjunction with feed agglomeration. Metal recovery at 85% gold and 75% silver is attained on properly agglomerated (pelletized) tailings that average 85% -75 µm (-200 mesh). The impounded tailings averaged more than 18% moisture (32% moisture maximum at depth) in place behind the storage dam. Therefore, the mining method, mixing and agglomeration technique, and materials handling practice were major considerations in making this operation a financial success. The Gooseberry Tailings Project is located about 35 km (22 miles) east of Sparks, NV. The project is owned by Asamera Minerals (US) Inc. The company's Western Nevada Gold prop¬erty surrounds the Gooseberry Mine by several miles on all sides. Gold and silver mineralization at the Gooseberry Mine is restricted to a narrow calcite-quartz vein. Calcite makes up more than 80% of the vein throughout the mine. The host rock consists of porphyritic andesites and dacites constituting volcanic flows. The rocks are propylitized or silicified to a varying degree. This ore mineralogy, then, also describes the tailings processed by heap leaching. The Gooseberry tailings were auger sampled in 1983 and again in 1987 after a final lift of 109 kt (120,000 st) was deposited in 1983 and 1984. The total volume behind the dam was calculated from planimeter measurements from tailings surface to original area topography.

Jan 1, 1990

By Ilpo O. Auranen

Background: Mines of today are moving ever increasing tonnages of rock. The new ore bodies tend be of lower grade and located deeper in the earth?s crust. The industry?s focus to overcome this ?deeper but poorer? ore problem has traditionally been to lower the costs by increasing the scale of mining. This has resulted in an ever increasing size of mining equipment like drills, loading shovels, dump trucks. The main focus has been in reducing the cost per ton of mining. This cost focus has often resulted in increased waste rock dilution and more ore losses. From costs to value: Waste rock dilution means waste rock in the concentrator feed. It?s value in open pit mining is equal to the value of ore it replaces. In underground mining, waste rock dilution?s value is equal to the value of ore it replaces added to the costs of mining it. It is commonly accepted that waste rock dilution in open pit mining ranges between 5 and 20% and in underground mining between 5 and 40 % . In a typical large open pit mine ore value is from 20 to 60 USD per ton and mining costs are 1 to 3 USD per ton. One percentage point reduction in waste rock dilution represents savings corresponding to about 20% of total mining operation costs. The difficulty of reducing waste rock dilution comes from the complicated nature of the ores. Accurate ore waste-border detection has not been possible due to excessive costs. Core drilling combined with geochemical lab analysis is the most accurate method for locating ore waste borders. However, it is also expensive and time consuming, which is why it is often combined with somewhat lower cost methods like sampling drill cuttings from blasting drill holes or from reverse circulation drills and analyzing them in a laboratory. The main drawback with these methods, in addition high costs, lies in the sampling accuracy and in the quality control of the whole process when a large number of samples have to be handled manually. In every mine it is possible to halve waste rock dilution by using new on-line analysis technology .

Oct 1, 2009

By Takashi Goto

In 1968 three historical rockbursts had occurred at Bibai Coal Mine in Hokkaido. At the time they had been observed by several seismic observatories as natural earthquakes away more than 100km from Bibai City. These facts taught us that the occuerrence of failure yielded in underground opening that was accompanied with seismic wave could be observed by seismometer on the surface. After two years from this desaster, the study concerning to microseismic events generated from underground mining panels has been begun by researchers with heart-throb.

Jan 1, 1982

This paper describes the oil engihe plant owned and operated by Gold Mines .of Kalgoorlie Limited, and gives some details of installation, operational, and maintenance experiences which have arisen during the years of service, the last 5 years of which have been under the writer's control. Nos. 1, 2, 3 and 4 alternator units have run 50,276, 48,006, 49,677 and 25,930 hours respectively, and much knowledge has been obtained of their performances and faults. During the war years, it has not been easy to obtain spare parts; this, coupled with the loss of No. 4: alternator unit to the Commonwealth Government, made operation and maintenance a very difficult problem.All engines are of the enclosed vertical, single acting, 4-stroke cycle, airless injection type.The electrical power plant originally consisted of four 6-cylinder Belliss & Morcom heavy oil engines, direct coupled to Bruce Peebles alternators, running at 375 r.p.m., and generating 3-phase electric power at 3,300 volts, 50 cycles.No. 1 unit, installed in 1936, is a 6-cylinder 450 b.h.p. standard 12A Belliss & Morcom heavy oil engine, direct coupled to 280 kw. 0.8 p.f. alternator. The remaining three units are 6-cylinder, 660 b.h.p., lSA oil cooled piston type Belliss & Morcom heavy oil engines, direct coupled to 410 kw. 0.8 p.f. alternators, making a total capacity of 1,510 kw.No. 4"alternator. unit, an 18A Bellis.& Morcom set, was taken over by the Commonwealth Government in June 1943.

Jan 1, 1945