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By J. H. Ambrose

Introduction The copper smelter of the Hudson Bay Mining and Smelting Company, Limited was constructed to handle the concentrates produced by the concentrator, the residues from the zinc plant, concentrates from Sherritt-Gordon Mines, Limited, and other custom material that might be available. Construction work was started in 1929 and the first concentrates were smelted in November, 1930. From that time until June, 1931, the operations were more or less of an experimental or tuning-up nature. In designing the smelter, it was believed the following tonnages would be available for smelting per day: Copper Tons Content H. B. M. & S. Co. concentrates 331 14.0 per cent H. B. M. & S. Co. zinc plant residues 77 2.5 per cent Sherritt-Gordon concentrates 200 20.0+ per cent Consideration was given to wet-smelting the charge, but, in order to give greater flexibility to the operation, it was finally decided to roast the concentrates before smelting. Briefly, the smelter equipment consists of three Nichols-Herreshoff roasters, one reverberatory furnace, two Peirce-Smith converters, a Walker casting wheel, and a four-unit Cottrell plant, two units' for the roaster gases and two for the converter gases. A general plan of the smelter is shown in Figure 1. Prior to the commencement of smelter operations, a stockpile of about 9,000 tons of concentrates had been built up by the concentrator and the old pilot-mill. The zinc plant was in operation, but the equipment for drying the residues at the smelter had not been completed. Sherritt-Gordon did not commence their shipments of concentrates until April, 1931. The only flux available was barren sand from a pit ten miles south of Flin Flon, adjacent to the railway. Several small deposits of metal-bearing siliceous ore had been discovered in the district, but none of these were sufficiently large to be considered as a source of smelter flux. At the beginning of March, 1931, the entire stockpile of concentrates had been consumed, and available tonnage for smelting was limited to the daily production of the concentrator.

Jan 1, 1935

By K Kongolo, M K. Kalenga

It is well known that the electrochemical precipitation of metals from aqueous solutions is conducted by the electrochemical standard potential scale. This scale indicates the order of nobility of different metals. Consequently, the standard potential scale shows the nocivity of metals when they are present as impurities in aqueous solutions containing metals. In the hydrometallurgy of cobalt, nickel and zinc are co-deposited during cobalt electrodeposition. Nickel has almost the same chemical, physical and electrochemical properties as cobalt. Furthermore, the electrochemical potential of nickel is relatively higher than that of cobalt. For this reason nickel is preferentially reduced at the cathode in the presence of cobalt. The presence of small amounts of zinc in solution leads to contamination of cobalt electro won. In spite of its lower electrochemical potential in comparison to cobalt, zinc is reduced on the cathode because of its higher hydrogen-overvoltage. But, in the hydrometallurgy of zinc, small amounts of cobalt present into solution will contaminates the zinc reduced at the cathode. This is easily explained by the fact that cobalt is more electropositive than zinc. The co-deposition mechanism of sulphur at the cathode is not yet well understood. Great attention has to be paid to the mechanism of co-deposition of sulphur during cobalt electrolysis..This work aimed to investigate the influence of some parameters on the co-deposition of nickel, zinc and sulphur during cobalt electrowinning. The following parameters have been investigated : Co2+ concentration in the electrolyte, current density and ammonium sulphate (NH4)2SO4 additions. It has been established that important nickel, zinc and sulphur co-deposition takes place at relatively low Co2+ concentration in solution. The behaviour of these elements with regard to the cathodic phenomena has been established.

Jan 1, 2005

By Henry M. Harris, Hal J. Kelly

Electrical dewatering tests are described in which the laboratory cells were designed for continuous operation using an endless metal belt as the anode . The test slurries were dilute , containing from 2.5 to 14.4 percent solids (specific gravity of 1.02 to 1.10 ) that were all minus 8 microns in size . Recovery of the solids of low water content was the practical objective of the investigation . Process variables evaluated and reported are : The effect of feed rate , amount of electrolyte , samples from different sources , and the use of two cells in series . These variables are evaluated on the basis of yield of clay per kilowatt - hour , recovery , and the amount of solids in the deposited clay . Recovery , yield , and solid content of the belt anode product are interdependent , and changes that will improve one of these characteristics often are accomplished at the expense of the others . Tests in which recoveries range from 50 to 60 percent usually result in yields of 8 to 9 kilograms (kg ) of dry clay per kilowatt- hour (kwhr ) . Under these conditions , the solid content of the deposited clay approximates 50 percent . One test sample , composed of kaolinite , showed very high yield , 13 to 16 kg per kwhr ; high recovery , 53 to 58 percent ; and 52 to 54 percent solids in the cake .

Jan 1, 1964

The coal-bearing areas of Kansas cover approximately 18,800 square miles (23 percent of the state's total area). Coal in Kansas ranges from lignite to high volatile bituminous, located principally in the south-eastern part of the state. Kansas reserves are estimated to be 987,500,000 tons, approximately 0.2 percent of U.S. reserves. Coal beds generally occur in thin seams up to five feet thick. The overburden-to-coal stripping ratios in Kansas are among the highest in the nation, typically 20 to 1, with some ratios up to 35 to 1. The first commercial coal mining operation in Kansas opened in 1865 at Fort Scott in Bourbon County. As in other states, coal production in Kansas increased during the industrial revolution, and then fell as industry converted to alternative fuels. Production peaked in 1916 and 1917, with annual production of 7,250,000 tons. Coal production today is by surface mining methods. The Kansas regulatory program was approved April 14, 1982; the Abandoned Mine Land Program was approved February 1, 1982; and the emergency program authority was approved January 10,1989.

Jan 1, 1992

By E. Asselin

Passivation of Fe-Ni-Co alloys during the Caron leach process is thought to result in low cobalt recoveries in ammoniacal ammonium carbonate solutions. A brief review of the available thermochemical data for the (Fe, Ni, Co)-NH3-H2O systems has been undertaken. The stability of the Co(II) and Co(III), Ni(II) and Fe(II) ammines was examined through the use of speciation and Pourbaix diagrams at 298 and 333 K and at 6 M total ammonia. Increasing the temperature reduces the stability fields of the ammines and generally constrains the thermodynamic window for the leach. The Fe(II) ammines are only thermodynamically stable under reducing conditions. Polarization experiments show that these ammines are unlikely to form if oxygen is present. Fe-Ni-Co alloy passivation may be alleviated by careful management of potential.

Nov 1, 2011

By E. Asselin

Passivation of Fe-Ni-Co alloys during the Caron leach process is thought to result in low cobalt recoveries in ammoniacal ammonium carbonate solutions. A brief review of the available thermochemical data for the (Fe, Ni, Co)-NH3-H2O systems has been undertaken. The stability of the Co(II) and Co(III), Ni(II) and Fe(II) am-mines was examined through the use of speciation and Pourbaix diagrams at 298 and 333 K and at 6 M total ammonia. Increasing the temperature reduces the stability fields of the ammines and generally constrains the thermodynamic window for the leach. The Fe(II) ammines are only thermodynamically stable under reducing conditions. Polarization experiments show that these ammines are unlikely to form if oxygen is present. Fe-Ni-Co alloy passivation may be alleviated by careful management of potential.

Jan 1, 2011

By Roy T. Bowles

This paper focusses on technological developments which occurred in the village of Deloro, Ontario between 1868 and 1919. These developments illustrate the manner in which improved metallurgical processes linked Ontario mineral resources to industrial expansion. Overview of Deloro The village of Deloro is in Hastings county, about 5 miles from the village of Marmora, the site of the old Marmora iron works and the more recent Marmoraton iron mine. In this area, the Canadian Shield meets the Great Lakes lowlands. The first settlement in the vicinity occurred in the 1820s and was associated with the Marmora iron works. Settlement grew around agriculture. Gold was discovered in Hastings county in 1868. Deloro - the name derives from gold - developed as a concentrated settlement in proximity to several gold mines and the related processing activities. By 1901, substantial smelting works were successfully producing gold and arsenic from local ores. The village was developing as a company town within the framework of benevolent paternalism. For reasons associated with markets and with local ore supplies, the Deloro mills were idle for a period beginning in 1902. Developments in later periods connected Deloro to world events. In 1903, silver was discovered at Cobalt, Ontario. By 1907, M.J. O'Brien, who owned mines at Cobalt, had purchased the Deloro works and was using them to process Cobalt-area ores*.

Jan 1, 1999

By J. A. Coll, R. W. Cahn, A. Lawley

WHILE the creep properties of pure face-centered-cubic and close-packed-hexagonal metals have been thoroughly investigated and are well established, body-centered-cubic metals have been studied less extensively. Moreover, very few fundamental studies on the creep of solid solutions, irrespective of crystal structure, have been reported. The present study is concerned with the creep of a series of body-centered-cubic solid solutions. The present position concerning creep of pure metals is, briefly, as follows.1"3 Creep at first takes place at a steadily decreasing rate; this is the stage termed primary or transient creep. Except at the lowest temperatures this is succeeded by a stage of secondary or steady-state creep. At high temperatures and stresses, this may be succeeded by an accelerating stage, termed tertiary creep, with which we shall not here be concerned. There is no well-defined physical model at present for the transient stage; in general terms, transient creep is best regarded simply as a manifestation of work-harden ing. Steady-stage creep can certainly take place by several different mechanisms: the choice of dominant mechanism depends primarily on temperature. We shall here be concerned only with high-temperature steady-state creep, a term usually reserved for creep at absolute temperatures higher than 0.5 Tm, where T, is the melting point. In this range, the activation energy for creep is, for many metals, equal to the activation energy for self-diffusion, and this is generally interpreted in terms of a "climb mechanism.1-4 The creep rate is determined by the speed at which dislocations, impeded by obstacles the nature of which is disputed but which are probably established during transient creep, can climb by means of a diffusion process, until they are able to by-pass the obstacle. In solid solutions, the intrinsic resistance to the slip motion of dislocations may be much larger than in the solvent, to the extent that the motion of dislocations in the glide plane, rather than their escape by climb out of this plane, may become the rate-controlling factor. weertman5 has considered this possibility from a theoretical point of view, and concluded that some form of "viscous slip" is likely to be rate-controlling at comparatively low stresses. The resistance to slip may arise from "atmospheres" of impurities forming around dislocations; a high Peierls force in materials of high cohesion; or some structural peculiarity such as clustering or ordering of solute atoms.= We shall be concerned here with the case of ordering. The only published investigations concerned explicitly with the effect of order on creep refer to creep in ß-brass by Herman and Brown,7 and in Ni-Fe alloys, by Kornilov and panasyuk8 and by Suzuki and Yamamoto.9 Recently, Herman and Brown's paper has been supplemented by a determination of the tensile yield point of ß-brass as a function of temperature.10 Both studies showed a sharp drop in resistance to deformation of ß-brass over a range of a few degrees just above the critical temperature Tc at which order finally disappears. These observations are especially noteworthy, because in ß brass the degree of order diminishes steadzly to zero as the temperature approaches Tc. It is, therefore, the disappearance of the last traces of long-range order which has the largest effect on the resistance of the alloy to plastic deformation. In the Ni/Fe alloys of various compositions, resistance to creep at a given temperature and stress is maximum at the stoichiometric composition, both below Tc, (long-range order), and above T,. (short-range order).' Near Tc, the creep resistance of an ordered alloy is much higher than that of the same alloy in the disordered condition.9 The aim of the present investigation was to study the creep behavior in the neighborhood of Tc, of another system of ordering alloys. The iron-aluminum alloys were considered the most suitable. because: i) The order again diminishes steadily to zero as the temperature approaches Tc; there is no sudden drop in order at Tc, and it therefore is

Jan 1, 1961

By Hannington MD

On the southern flank of Lihir Island, a group of three volcanic cones were discovered at water depths from 1000-1500 m. The volcanoes are located in a narrow zone of recent seismic activity and elevated heat flow (up to 100 mW/m2). The recovered rocks consist of fresh alkali-olivine basalts, clinopyroxene-rich basalts, and porphyritic phlogopite basalts. One of the volcanoes has a recently erupted ejecta blanket (recovery of warm ash!), containing a suite of abundant mafic to ultramafic xenoliths. These xenoliths appear to represent a unique cross-section of subduction- modified mantle material sampled by volatile-rich primary fore-arc magma during its ascent to the seafloor. The most spectacular finding was the discovery of an active hydrothermal system with associated vent fauna and mineralization at Edison Seamount, and the recovery of unusually gold-rich (up to 43 ppm Au) hydrothermal precipitates in 1050 m water depth at the top of Conical Seamount, only 25 km south of the world-class epithermal Ladolam gold deposit on Lihir. The discovery of submarine hydrothermal vents associated with highly alkaline, quartz-undersaturated fore-arc volcanics represents a significant departure from similar hydrothermal systems on the mid-ocean ridges and in back-arc basins. The high gold grades in hydrothermal precipitates at Conical Seamount may indicate the first documented example of a shallow-marine epithermal system analogous to those known on land, and has important implications for the gold potential of submarine arc environments.

Jan 1, 1995

By Sean Gallagher, Thomas G. Bobick

The Bureau of Mines performed a pilot study examining the effects of posture on back strength and Maximum Acceptable Weight of Lift (MAWL) on six healthy male subjects (M = 32 years + 4 SD). Six hack strength measurements (3 static and 3 dynamic) were made while the subjects were kneeling and standing. In addition, these subjects (who were unaccustomed to lifting in these postures) volunteered to participate in a study of psychophysically determined NAWL in both postures. Results of the back strength tests showed a significantly lower peak torque per body weight output in kneeling versus standing hack strength measurements for five out of six test comparisons (p < .05). Subjective estimates of lifting capacity in the kneeling posture were significantly lower than those for the stooped posture (p < .05). The results of tests of hack strength and lifting capacity in these two postures provide useful information to consider in determining the physiological and psychophysical stresses imposed by these

By Doss H. White

In 1989, Alabama&apos;s nonfuel mineral industry produced a State record of $463 million of nonfuel mineral commodities, an increase of $4 million over that reported by the State&apos;s mineral producers in 1988. The new record marked the fourth consecutive year in which mineral production increased over that of the previous year. Alabama ranked 22d among the States in total mineral value. Cement, stone, and sand and gravel were the three leading mineral commodities produced, accounting for 77% of Alabama&apos;s total mineral value. During the year, there were several acquisitions by Alabama cement and stone companies, and the only tantalum mine in the United States completed its first full year of operation.

Jan 1, 1991

By F. von Gemmingen, E. J. Fasiska, L. Zwell

EARLIER studies of the dependence of creep on structure1 and of the dependence of creep rupture on creep behavior2 did not include detailed discussion of the phases present in the specimens under examination. It is of interest to determine the nature and amounts of such phases present since it is well known that these are dependent on composition, heat treatment and creep conditions.3 This investigation has therefore been undertaken to identify these constituents and to determine their relative amounts. This report presents the results of this investigation and describes an improved "extraction replica" technique developed in the course of the work. The composition of the Type 316 austenitic stainless steels studied in this investigation is given in Table I. The specimens were solution treated at 2000°F for 1/2 hr and water quenched. Those specimens subjected to prestrain at room temperature Methods for extracting precipitate particles from materials for electron and X-ray diffraction analysis have been described by Fisher,4 Smith and Nutting: and Booker and coworkers.6&apos;7 Electrolytic extraction methods have been described by Andrews and Hughes.8 The first two methods have been used at this Laboratory, but have not proved satisfactory for many problems requiring X-ray diffraction analysis. It has been difficult to remove the replicas or to avoid tearing them during removal. These difficulties are overcome by backing the replica with a water-soluble plastic such as Dupont 51-05 Elvanol (polyvinyl alcohol). The backing increases the thickness of the replica so that it can be mechanically stripped from unloosened particles without folding or tearing. This procedure eliminates the stripping of wet replicas, which usually have a low tear strength. By stripping replicas after drying, it is possible to extract partially loosened particles which otherwise would remain imbedded in the metal. The modified method is as follows: after the specimen is polished, it is etched for 60 min in modified glyceregia consisting of 30 ml glycerol, 7 ml nitric acid, and 4 ml of hydrochloric acid. The sample is then washed, air dried, and coated with a 20 pct Baker and Adamson 1611 collodion (85 pct amyl acetate by volume) film. After the collodion film is dry (about 15 min), the sample is again etched in the modified glyceregia for 50 min, washed with isopropyl alcohol (to prevent dissolving the collodion) and air dried. A very thick coat of polyvinyl alcohol (about 50 pct solution "Elvanol Dupont 51-05" in water) is poured over the sample and is permitted to dry for about 24 hr. The composite film containing the precipitates is then stripped from the specimen and floated, polyvinyl alcohol side down, in water (at about 50°C) until the polyvinyl alcohol dissolves. This usually takes about 15 min. Should any of the polyvinyl alcohol remain on the replica, it does not give a crystalline diffraction pattern and therefore does not hinder identification of the precipitates. The extraction replica, floating in water, is picked up on a glass fiber and mounted in the Debye-Scherrer camera for X-ray diffraction analysis. Concentration of the etchant, and also etching time, had to be increased for several of the samples. Fig. 1 shows two extraction replicas taken from different areas of spec. M74-37; X-ray analysis showed a strong diffraction pattern of M23C6 and a medium one of u phase. A

Jan 1, 1962

By Alfredo Bambaren

First Encounters between Communities and Mining 1. Importance, meaning and process 2. Antamina?s study case 3. ?Self-organization? theory (NLP) 4. The story we tell, the values we live 1. Importance, Meaning and Process ? First encounter stays in mind as a model. ? First encounter means foundation. ? First encounter builds the story.

May 1, 2010

By I. R. Firth, D. L. Taylor

The ability to accurately model the movement of grade boundaries during a typical bench blast would be of significant benefit to mine operators. The movement of grade boundaries during the blasting process may have significant impact on grade recovery if not accounted for. The paper describes the preliminary stages of a research project which combines numerical blast models and conventional mine planning software to allow practical use of the results of a numerical blast simulation. A two dimensional bench blast is simulated using the Universal Distinct Element Code (UDEC). The results of the simulation are used to provide input to the Surpac mine planning package which can be used to generate grade control plans.

Jan 1, 2001

By Iga Trisnawati, Fitria Puspita, Himawan Tri Bayu Murti Petrus, Gyan Prameswara, Idi Amin, Andi Nurfaadhilah Ulfah

Laterite ores, rich in essential metals like nickel, iron and cobalt, are a crucial resource for modern technology. However, the extraction of nickel from laterite ores is a complex process due to their intricate mineral compositions and low nickel concentrations. This paper presents the results of a study to understand the leaching behavior of preroasted laterite ore and the impact of roasting on leaching kinetics. The ore was subjected to roasting at 280 and 610 °C for a duration of three hours, followed by leaching in sulfuric acid (H2SO4) under atmospheric conditions. The results show that preroasting the ore at 610 °C significantly improved the recovery of nickel during leaching, with recovery rate of up to 93 percent, and the leaching process of the preroasted ore was primarily controlled by diffusion through the ash layer. The activation energy values were found to be 20.97 kJ/mol for the ore roasted at 280 °C, and 18.97 kJ/mol for the ore roasted at 610°C. These findings suggest that preroasting can enhance the leaching performance of laterite ore, potentially paving the way for more efficient and sustainable nickel extraction processes.

Aug 1, 2024

By E. G. Leonardon, C. Schlumberger, M. Schlumberger

Last year the authors presented a paper that discussed the various electrical measurements they perform in drill holes, which they name "electrical coring."&apos; The object of the present paper is to give a brief account of the points of scientific interest that have been elucidated by further field work and laboratory research, as well as the economic results that have been attained recently. The practical examples given in this paper are drawn, for the most part, from the field work performed in Russian oil fields during the years 1931 and 1932. This is due to the fact that, thanks to the comprehensive cooperation of the Russian geologists and scientists, the authors are allowed to make public a large amount of recant information. In other countries in which the authors are engaged in similar work, the competition between private companies makes it imperative, for a considerable lapse of time, to avoid publication of the results obtained. The following points will be discussed successively, under separate headings: 1. Electrochemical phenomena in drill holes. 2. Stratigraphic correlations and tectonic studies. 3. Relations&apos; between the productivity of oil horizons and their electrical resistivity. 4. Location of coal seams. 5. Summary and conclusions. Electrochemical Phenomena in Drill Holes Technical Publication 462l described the author&apos;s discovery concerning the existence of important differences of potential spontaneously gener-

Jan 1, 1934

By W. W. Adams

With production of coal per man-hour of work at a higher level than ever before, the coal-mining industry of the United States established a near-record in lowering the accident rate of nonfatal injuries per million man-hours of exposure in 1940, the injury rate being lower, except for 1939, than in any year since 1930, the first year for which comparable figures for nonfatal accidents were collected by the. Bureau of Mines, United States Department of the Interior. The rate for fatal accidents, on the other hand, was not as low or favorable as in 1939, nor was it as low as in many years since 1910, the earliest year for which fatality rates were complete for all coal-mining States. Accident and employment reports received from mine operators and fatal-accident reports from State mining departments and inspectors showed that 533,267 men were employed at coal mules in 1940, exclusive of the number employed at coke ovens, Although this number represented 6,108 fewer employees than in 1939, a large increase was made in the volume of work performed and in the quan¬tity of coal produced. In 1940 the amount of labor performed at coal mines was 748 million man-hours, a gain of more than 10 percent over the previous year. The average employee worked 199 days, an increase of 20 workdays per man. Production of coal was nearly 513 million tons, an increase of 14 percent, of which more than 461 million tons was bituminous coal and more than 51 million tons was Pennsylvania anthracite. Accidents to mine workers caused the death of 1,388 employees; in addition, 59,781 employees were injured, and each injury resulted in disability of the employee for more than the remainder of the day on which the accident occurred.

Jan 1, 1942

By Arthur. H. Winckers, Stephan L. Chryssoulis

"The deportment of gold in residue samples of cyanide leach tests of David Bell ore from level 6A block 13 was established using a comprehensive quantitative mineralogical approach. In every test the feed was first pre-aerated for 4 hours, followed by 48h cyanidation. Test no. 2 was done without lead nitrate and yielded a residue assaying 3.03g Au/t. In test nos. 4 and 9 lead nitrate was added before pre-aeration at 0.03kg/t and 0.3kg/t respectively yielding residues of 1.04 and 0.81 g Au/t.The mineralogical approach employed to establish the deportment of gold involved superpanning, to preconcentrate the free gold and assaying of the tailing fraction to determine the gold associated with quartz. It was thus determined that without lead nitrate added 47.2% of the lost gold is in the form of residual free gold particles less than 20 µm in size. It is this form of gold which is under represented in the residues of leach tests with lead nitrate added.Having eliminated coarseness of gold particles as a factor for lower recovery, it was determined that the residual gold particles have extensively contaminated surfaces. Amongst the deleterious surface contaminants identified: S, As, Ag, Sb and Hg, the two most harmful are S and Hg because of their concentration on the surface and rate of accumulation.Mercury and Ag are minor constituents of native gold at David Bell (average composition Au: 78.9 wt%, Hg: 18.5 wt% and Ag: 2.6 wt%) and they become enriched on the surface of residual gold particles because of their slower rate of dissolution in cyanide. The other elements absorb onto the gold particles from solution.The beneficial effect of lead nitrate was documented by comparative surface analysis of gold from residue samples of 48 hour leach testing with and without lead nitrate added and from time dependent (0-2 hour) leach tests in the presence and absence of lead nitrate. Lead nitrate accelerates the dissolution of gold by performing the following functions: (i) catalytic effect on the cyanidation of gold by accepting electrons; (ii) reaction with surface sulphur to form PbS which is then removed by oxidation; (iii) catalytic reaction on amalgam and Ag facilitating the cyanidation of Hg and Ag thus slowing down their build up onto the dissolving gold particles and (iv) the precipitation of sulphide sulphur from solution thus minimizing its build up on the surface of dissolving gold particles which is believed to facilitate the access of cyanide ions."

Jan 1, 1996

By Shunhong Huang

Thirty-five soils were sampled from three sites at different distances from chromium slag heap of a Iron-alloy factory, namely, nine from site A located under the chromium slag and eighteen from Site B in the vicinity of the factory and eight from site C far from the factory. The contents of total Cr in these soils from site A, B and C were 1666.0±1310.9 mg? kg - 1,1136.2±1844.6 mg? kg - 1 ,300.4 ±165.1 mg? kg ? 1, respectively. The contents of water soluble Cr were 59.3±77.57 mg? kg ? 1,0.6 ±0.3 mg? kg - 1 ,0.5±0.3 mg? kg - 1 , respectively. In order to evaluate soil health, the activities of four soil enzymes (catalase, polyphenol oxidase, dehydrogenase, alkaline phosphatases) were analyzed. The results have shown that dehydrogenase activity was greatly depressed by conditions in the heavy chromium-contaminated sites. Negative correlation was observed between dehydrogenase activity and the distance from the chromium slag heap, however, the activities of catalase, polyphenol oxidase, alkaline phosphatases in soils from the three sampling sites mentioned above did not differ significantly. The result imply that of dehydrogenase activity could be used as an indicator for the chromium pollution level in the area of Iron-Alloy factory and for the stability level of the soil ecosystem.

Jan 1, 2008

By Don Macfarlane, Brian Heer, Tom Martin

The Manapouri Power Station, commisioned in 1969, is located in Fiordland National Park in the South Island of New Zealand. It is the country’s largest hydro power scheme and can generate 700MW of electricity. The power station utilizes the hydraulic head of Lake Manapouri, using an underground power station and a 10 km-long tailrace tunnel was constructed by drill-and-blast to allow discharge to Doubtful Sound. Power generation fell short of expectations due to higher than expected friction in the tailrace system. To remedy this a second tailrace tunnel was constructed between June 1998and May 2002. It consisted of a 10 m diameter, 10 km long tunnel in predominately gneissic rock and was excavated using an Atlas Copco Robbins TBM. High water inflows, Owner/Contractor production incentives and a major claim settlement were handled during the 5 1/4 year project. This paper discusses TBM advance rates, utilization factors that increased in the second half of the project plus an innovative method for sealing Tunnel 1 to construct the bifurcation.

Jan 1, 2003