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By Herbert H. Einstein

Swelling ground is one of the most feared problem in tunneling. Although usually not causing sudden failures during construction, the very large and long- lasting deformations give contractors and designers the impression of facing an inevitable phenomenon against which there is no technical defense. Examples are concrete liners of up to 1 m that have been cracked or crushed, tunnel inverts that are heaving at a rate of 0.1 m per week during construction and continue to heave during decades. This author (Einstein (1979). Einstein et al. (1972) (1975)) and others (Huller (19761, Grob (1972) (1976). Golta (1967), tienke (19761, Krause (1976). Lee et al. (1976)) have described many cases of swelling tunnels that illustrate the significance and magnitude of the problem. This paper intends to show that tunnels in swelling rock can be built to perform satisfactorily. . For this purpose the tunnel builder needs to know how to determine the existence and specific behavior of swelling ground, how to select appropriate structural characteristics of the tunnel and predict its performance under the specific ground conditions, and finally how to construct the tunnel such that the anticipated performance will be achieved. This concept of a design-construction procedure is the same as that for tunneling in general; it is mentioned here for NO reasons: 1. To point out that design and construction cannot and should not be separated. The way in which the tunnel will be built has to be an integral part of the design. 2. This paper will be structured corresponding to the procedural concept.

Jan 1, 1981

By J. B. Yianatos, F. A. Contreras

Mineral carrying rates across the pulp-froth interface in the range of 1.4-3.2 tph/m2 and bubble surface coverages between 7 and 22 %, have been determined in large rougher flotation cells operated under normal conditions. To evaluate these variables, measurements of bubble size distribution, bubble load and size distribution of particles collected by true flotation, were performed at different concentrators. A model for estimating the bubble surface coverage at the pulp-froth interface level as a function of bubble load, Sauter mean bubble diameter and particle size distribution of the mineral collected by true flotation, was developed. Bubble loadings in the range 24-70 kg/m3, bubble sizes between 1.5 and 3.2 mm and superficial gas rates between 1.2 and 2.0 emfs, were measured in copper rougher flotation cells under normal operating conditions. Estimated bubble surface area fluxes were in the range 31 to 62 cm2/s/cm2- Results of carrying rates and bubble surface coverage, at the pulp-froth interface, were significantly lower than the maximum carrying rate for the collection zone, indicating that in this case is difficult to achieve maximum carrying rate at the pulp-froth interface level. The main constraints regarding carrying capacity limitations are then related to the froth transport characteristics (i.e. cell design, froth residence time, froth stability, froth recovery). Advances in Mineral Processing Science and Technology Proceedings of the 48th Annual Conference of Metallurgists ofCIM Sudbury, Ontario, Canada Edited by

Jan 1, 2009

By W. B. Boggs

A paper describing the Noranda smelter was presented at the 1930 annual meeting of the Institute (1). A lacer article appeared in the Trans-actions of the A.I.M.&M.E. in 1933. Since that time there have been many changes and improvements in plant operation and it is the abject of this paper to give an account of developments up to the beginning of 1939. The Noranda smelter was built originally for a capacity of 30,000 tons a month. At present, the average tonnage being treated is 110,000 tons a month, the tonnage of custom ores alone amounting to almost the total rated capacity of the original plant. This increase has been accomplished by increasing the capacity of the reverberatory furnaces from 700 tons of solid charge per furnace day in 1928 to 1,500 tons per day at the present time. In 1935, it appeared chat the limit of furnace capacity had been reached at about 1,300 tons of charge per day because of the short life of the sprung silica arches. However, the adoption of the suspended magnesite roof, which had been developed at the Flin Flou smelter of the Hudson Bay Mining and Smelting Company, overcame this difficulty and it is believed that the limit of the capacity of the present reverberarory furnaces has not been reached. A brief account of the changes to equipment in the various departments will be given, followed by a description of the more important features of present smelter practice.

Jan 1, 1939

By Raffaella Granata, Daniele Vanni

"This paper presents the drilling and grouting program performed at two abandoned landfills located into former calcarenite rock quarries in Manfredonia, Italy. Tests performed after the abandonment of the landfills detected leachate infiltration into the groundwater, which required the implementation of a remedial program. In order to seal the rock fissures and create an impervious basin at the base of the two landfills, sub-horizontal boreholes were drilled and grouted into the rock under the waste deposits. At the same time, sub-vertical boreholes were drilled and grouted along the perimeter, to laterally confine the landfills. Finally, the landfills were properly capped with an HDPE membrane, geogrid, geotextile and topsoil. The methodology implemented to create a watertight barrier under the two landfills was innovative and cost-effective. Moreover, the drilling and grouting minimized the environmental impact of the remedial work, since the waste remained in their original location and were not affected by the remedial process.IntroductionThe two landfills, “Pariti-1 RSU” and “Conte di Troia”, are located approximately 6 km (4 miles) southwest of Manfredonia, a small town close to Foggia, Italy (Fig. 1). These landfills are located into two rock quarries, formerly used for the extraction of calcarenite (i.e., a weak calcareous sandstone, in tuff facies), and were abandoned in the early 90s. The two landfills cover a total area of about 45,000 m2 (11.2 acres), and mainly include urban solid waste with a depth of up to 25 m (82 ft). The main characteristics of the two waste landfills are listed in Table 1."

Jan 1, 2016

By G R. Hancock, M J. Saynor, J B. C Lowry, K G. Evans, M Narayan

Tools such as landform evolution models (LEMs) provide a means of predicting how a landform may evolve over extended periods of time. Here, we utilise the CAESAR-Lisflood LEM to assess a proposed rehabilitated landform of the Ranger Uranium Mine in the Northern Territory of Australia (Figure 1). The environmental requirements for the closure of the mine state that the final landform must be able to physically isolate tailings material from the environment for a period of at least 10,000 years (Australian Government, 1999). However, uncertainty exists as to the reliability of model predictions representing thousands of years.In this study, we use model predictions from an undisturbed pre-mine landform to calibrate the parameters used in model simulations of a post-mining rehabilitated landform over a simulated period of 1,000 years. Specifically, we vary environmental conditions such as rainfall, presence/absence of vegetation and surface particle size characteristics in a series of model simulations to determine their impact on the respective landforms, to revise model input parameters as required. A simulated period of 1,000 years was used as it represents a period in which long term landform evolution patterns and trends may be established and identified. CITATION:Lowry, J B C, Narayan, M, Evans, K G, Saynor, M J and Hancock, G R, 2018. Using an undisturbed landform to calibrate long-term predictions of the evolution of a rehabilitated landform, in Proceedings Life-of-Mine 2018, pp 133–137 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 25, 2018

By Smith N. S, Rholl S. A, Otterness R. E

The Bureau of Mines is conducting research to develop blast designs for the control of fragmentation and rock displacement and the reduction of overall mining costs. This Bureau experiment, a program of 25 reduced-scale blasts conducted at the University of Missouri-Rolla's experimental quarry, determined the effects on fragmentation of changes in burden, spacing, explosive diameter, and stemming-explosive length. Combined with previous work at this experimental quarry examining air gap, delay, and decoupling effects, this experiment provided a data base of over 50 shots. The fragmented rock, 1000 to 3000 kg from each 3- to 4-hole blast in the 1 m bench of massive dolomite, was completely screened to assess fragmentation. The explosive used for all the tests was extra dynamite initiated by seismic caps fired from a sequential blasting machine with delays of 5.2- to 13-ms/m of burden. Empirical equations were developed which predicted the effect of shot design parameters on the fragmentation. These formulas predicted the average fragment size of each of the 25-shots to within 13 percent. For spacing to burden ratios of 1, 1.4 and 2, there were optimum burden to explosive diameter ratios (B/de) of 30, 26 and 20, respectively. Decreasing the B/d. ratio beyond these levels resulted in little or no reduction in the average size. For B/de ratios above these levels the average size increased but, at least a 15 percent reduction in the average size results when the shot design parameters are optimized.

Jan 1, 1990

By Ryan A, Dennis R, Dekker D, Ellis W, Sowerby B

An automated facility for non-destructive elemental analysis of mine samples has been developed and tested at Mount Isa in association with the CSIRO. The analysis is based on the prompt and delayed measurement of gamma radiation emitted by the sample during and after separate fast and thermal neutron irradiation. No sub-sampling or sample preparation is required and the analyser is operated by one person that would otherwise be splitting drill core in preparation for conventional analysis. The analyser has been calibrated to measure the following (precision in parenthesis): Ag (14 ppm), Pb (0.5 wt%), Zn (0.5 wt%), Cu (0.04 wt%), Fe (0.7 wt%), S (1.1 wt%), SiO2 (0.8 wt%), A1203 (0.07 wt%), Mn (0.05 wt%), specific gravity (0.02 g cmt). Crushed or whole drill core samples that range in size between 0.5 m and 1.8 m of equivalent BQ size drill core can be analysed at a rate of six samples per hour. If required, drill core samples can be re-packed in their original configuration for storage following analysis. Standard samples are measured every day to ensure operational stability. The analyser is controlled by a personal computer which prints the results and stores them along with the raw data on the computer hard disk. This paper reports the performance of the analyser during 1992 for routine analysis of drill core from the lead ore bodies of both Mount Isa and Hilton mines.

Jan 1, 1993

By Charles Y. Clayton

A study of the literature shows that the greater part of research work on annealing of cold-worked iron has been for the purpose of studying the effect on grain-size and properties other than hardness. No reference has been found of experimental work of the same nature as that explained in this short paper. The material used throughout was Vismera iron, an iron containing 0.03 carbon and made several years ago by the Inland Steel Co. One-half inch stock was cut into cylinders % in. long. The cylinders were then compressed for 60 sec. in a Reihle testing machine under loads as indicated below: CoMPressiON, Average Length, Series Pounds Inches 1 10,000 0.693 2 15,000 .599 3 20,000 .501 4 25,000 .426 5 30,000 .378 6 35,000 .351 7 40,000 .311 Fig. 1 shows these cylinders after compression. Each series consisted of 17 specimens lettered 0, B, C, D, E, F, G, H, I, J, K, L, M, N, P, Q and R. One specimen of each series was reserved for study of the properties in the cold-worked condition. The other specimens were annealed in a Hump furnace for 1/2 hr. at the following temperatures: TEmpeRa- TempeRa- Letters tuRe,° C. LettERs tuRe °C. B............................... 250 J.............................. 650 C............................... 300 K............................. 700 D............................... 350 L.............................. 750 E............................... 400 M............................. 800 F................................ 450 N............................. 851 G............................... 500 0. . :........................... 900 H............................... 550 P.............................. 950 I................................ 600 Q............................. L000

Jan 1, 1926

By M Nagarajan, S Esen

This paper presents three case studies from three surface coalmines. The first case study is from a coalmine where a dragline operates in 60 m wide and 1 km long strips. Typical overburden height ranges from 30 m to 50 m. The base case blast was modelled using an advanced blasting model (Distinct Motion Code – DMC) to establish the baseline using the blasting parameters in use before optimisation. DMC was used to model the alternative blasting scenarios to obtain the following key objectives:the post-blast muck pile profile was required to have an optimum height of ~28 moptimise cast percentageachieve a well-fragmented muck pile for the draglinethe muck pile profile should be appropriate for the dragline entry into the strip.Achieving these objectives would optimise the dragline performance through reducing rehandle and improving advance along the strip. DMC modelling resulted in the following outcomes:the required muck pile profile could be achievedthe cast percentage increased from 21.1 to 25.1 – a four per cent increaserehandle volumes were reduced.Establishing a good quality assurance / quality control (QA/QC) process was essential to delivering a blast that could achieve these outcomes. The blast was fired successfully and the dragline completed the strip approximately two weeks ahead of the time taken for the previous strip.The second case study is from a coalmine where D11 dozers are used to move the blasted material to the spoil. An optimum profile was required for the dozers for this low-cost mining operation. To achieve this, an electronic blasting system (uni tronicTM 600) was introduced to optimise the muck pile profiles being delivered. The new muck pile profiles showed improved cast percentage from 10.6 per cent to 17.8 per cent. The changes in the muck pile centre of mass displacement improved dozer efficiency as a result of pushing material shorter distances to its final position. Dozer productivity (bcm/h) statistically did not change but coal was being exposed at a faster rate.The third case study summarises the results from a site where dozers are used to minimise the total mining cost. Traditionally, excavators had been used on-site with conventional blast designs. After firing the first cast shot, the blasted material was found reasonably well-fragmented. Although dozing productivity hasn’t been quantified on-site, operators were providing positive feedback (easier push) due to the reasonably good muck pile profile suiting to the dozers. This site also realised significant savings with the dozing as compared to their truck/shovel operation.CITATION:Esen, S and Nagarajan, M, 2015. Muck pile shaping for draglines and dozers at surface coalmines, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 409–416 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By W. R. Reed, G. J. Joy, B. Kendall, A. Bailey

"Testing of the roof bolter canopy air curtain (CAC) designed by the U.S. National Institute for Occupational Safety and Health (NIOSH) has gone through many iterations, demonstrating successful dust control performance under controlled laboratory conditions. J.H. Fletcher & Co., an original equipment manufacturer of mining equipment, further developed the concept by incorporating it into the design of its roof bolting machines. In the present work, laboratory testing was conducted, showing dust control efficiencies ranging from 17.2 to 24.5 percent. Subsequent computational fluid dynamics (CFD) analysis revealed limitations in the design, and a potential improvement was analyzed and recommended. As a result, a new CAC design is being developed, incorporating the results of the testing and CFD analysis.IntroductionExposure to respirable coal mine dust can cause coal workers’ pneumo-coniosis (CWP), also known as black lung. If silica is present in the dust, the miner may develop silicosis in addition to CWP. CWP and silicosis are occupational respiratory diseases that are ultimately fatal and have no cure. The only method of prevention is to eliminate exposure to respirable coal mine dust and crystalline silica, or quartz. The occupational exposure limit on a Mining Research Establishment (MRE) basis for respirable coal mine dust had been 2 mg/m3 during each shift that a miner is exposed in the active workings of the mine or in mine facilities. Under recently enacted legislation, this standard was changed in August 2016 to 1.5 mg/m3 for a full working shift (The Office of the Federal Register, 30 CFR 70.100, 2015). When respirable quartz is present, the mine must maintain its average concentration at or below 0.1 mg/m3. If the mine exceeds the 0.1 mg/m3 respirable quartz dust concentration, then the applicable respirable dust standard is reduced, calculated as 10 divided by the percentage of quartz present (The Office of the Federal Register, 30 CFR 70.101, 2015).The sources of exposure to respirable coal mine dust for roof bolting occupations include infrequent maintenance and cleaning of the vacuum dust collection system and operating the bolter downwind — in the return air — of the continuous miner (Good-man and Organiscak, 2003). Recent studies have demonstrated that vacu-um dust collection systems are effective in capturing bolter-generated dust if maintained and operated correctly and if proper face ventilation techniques are used. However, most expo-sure to respirable coal mine dust for a roof bolter occurs when the roof bolter machine is operated downwind of the continuous miner (Potts, Reed and Colinet, 2011; Colinet, Reed and Potts, 2013). Because of this higher exposure downwind of the continuous miner, the roof bolter is often limited to working downwind of the miner only once during a production shift."

Jan 1, 2017

By D. T. A. Symons, M. M. Kapnistos

"Gravity profiles were established over five known Middle Silurian pinnacle reef s in Lambton County, southwestern Ontario. The profiles were approximately 3 km long, with stations every 60 m. Four fairly recent technological developments were used which now make the gravity method relatively efficient and inexpensive compared to the techniques available two decades ago. These are: (1) a La Coste-Romberg model G gravimeter to measure the gravitational attraction with a sensitivity of 0.01 mgal; (2) a Sokkisha model B-2 semiautomatic engineers level to measure the station elevations with an accuracy of 0.5 cm; (3) a Huntec FS-3 hammer reflection seismograph to determine the overburden thickness to an accuracy of ± 7%; and (4) a modern microcomputer to process the data sets and mathematically filter the gravity values. Thus, two men can measure and analyze approximately 1 km of profile per day and obtain the residual gravity profile to an accuracy of ± 0. 05 mgal. Positive gravity anomalies of 0.17 to 0.72 mgal were observed along all five profiles that approximately delineate the dimensions of the known pinnacle reefs, which have an average height of 120 m, diameter of 1500 m and depth to top of 575 m. The relative ease of the gravity survey method makes it a cost-effective technique for hydrocarbon exploration in southwestern Ontario. IntroductionPinnacle reefs are important sources of oil and gas. Although the pinnacle reefs in southwestern Ontario are small, they may contain commercial pools of oil and gas because of their proximity to refineries. They are also useful for subsurface storage of gas to meet the winter surge in gas demand.At present, geophysical exploration for new pinnacle reefs is done primarily by relatively expensive seismic methods, which are also limited in southwestern Ontario by prohibitions on blasting. Gravity methods of exploration were attempted more than two decades ago in this region. (Pohly, 1954); however, the level of technology at that time made the method marginal in sensitivity and quite expensive. Several subsequent developments- gravimeters with greater sensitivity and operating ease, ""self-levelling"" levels for elevation control, hammer seismic reflection units for determining overburden thickness and portable microcomputers for efficient data analysis-have significantly improved the precision and cost effectiveness of the gravity method. Five known pinnacle reefs were selected in Lambton County, southwestern Ontario (Fig. 1), to test the effectiveness of integrating these developments to locate and define additional reefs. Isolated reefs were chosen to avoid the problem of the gravity effect s of one reef interfering with those of a neighbouring reef. The surveys were facilitated by choosing reefs with a concession road passing directly over them ."

Jan 1, 1980

By S. O. Castro, R. E. González, L. E. Velasco, J. M. Van Sint, P. V. Lois

Abstract Several zones with expansive rocks were encountered during construction of two tunnels of the water conduction system for the Chacabuquito Hydropower Plant, located in the Aconcagua river valley of the Chilean Central Andes Mountains, at elevation 1.075 m a.s.l., and 65 km north-northeast of Santiago. The expansive behavior was associated to three different geological features. Firstly with the crossing of important north-south trending faults, with montmorillonitic and caolinitic fillings. Secondly to the presence of zones of rock mass showing a penetrative presence of clay minerals (montmorillonite, illite and caolinite) and penetrative thin veins of ceolites. And thirdly to the hydrothermal alteration of the rock and the consequent generation of clay minerals (predominantly montmorillonite, sericite and minor caolinite). The geological and mechanical characteristics of the expansive rock domains, as well as the field and laboratory reconnaissance and characterization program, are described in the paper together with the bases for design and the guidelines for construction and definitive tunnels support. Resúmè Pendant l’excavation des deux tunels du système d’adduction d’eau pour alimenter la centrale hidroélectrique Chacabuquito, plusieures zones de rochers expansifs ont été découvertes. Ces tunnels sont dans la vallée du fleuve Aconcagua de la Cordillère des Andes du Chili Central, à une altitude aproximée de 1.075m s.m.n. et à 65km vers le nord-nord-ouest de Santiago. Cette conduite expansive est associée à trois différentes caractéristiques géologiques du massif rocheux. Un premier lieu, à l’intersection de plusieures zones de failles, d’orientation nord-sud et subverticales, avec remplissage de rochers désintégrés, montmorillonite, caolinite et sérite. En seconde lieu, a la présence de zones de rochers expansifs, avec des nombreux minces veines de céolite-calcite et aussi des minéraux argileux (montmorillonite, illite et caolinite). Et, dans troisième lieu, dût à la présence de zones d’altération hidrothermale du massif rocheux et la conséquente génération de minéraux de remplacement argileux (surtout sérite, montmorillonite et en plus petit degré, caolinite). Les caractéristiques géologiques et géomécaniques des rochers expansifs, ainsi que la reconnaissance du terrain et les résultats du programme d’essais de caractérisation géomécanique du laboratoire sont décrits dans ce travail, ainsi que les bases de dessin et recommandations constructives pour placement du support définitifs pour ces tunnels Zusammenfassung Während der ausgrabung der zwei tunnels der wasserleitungssystem, um die Chacabuquito Wasserstromzentrale zu besorgen, wurden verschieden ausgedehnte gesteinadern gefunden Die beiden tunnels befinden sich im Aconcaguatal, wo der Aconcagua Fluss fliesst, auf einer Höhe von 1.075 m über dem Wasserspiegel und ca. 65 Km Nord-Nordest von Santiago entfernt, in mittleren Anden Cordillere (chilenische Gebirgskette). Drei verschiedene geologische merkmale dieser massiven gesteine können dieses ausgedehnte verhalten erklären. Erstens, hier befinden sich verschiedene und auch ganz wichtige

Jan 1, 2003

By N. Arbiter

During the first four decades of flotation practice, close to 200 machine designs were patented, although only about a score were ever used commercially. By the end of that period, the rigors of use had resulted in obsolescence of most pneumatic cells, with a few mechanical cells dominating the field. After World War II, newer designs began to appear but most were modifications in varying degrees of older designs with special features added. Of the dozen or so most recent developments, almost all are pneumatic, relying on compressed air instead of impellers for motive power, although in several cases pumps supplement compressors. Table 1 lists a few of the more fully documented new cells, with their significant characteristics, along with similar characteristics of a widely used mechanical cell. New designs in general are emphasizing: bubble size control; in several cases, methods for delivering feed directly to the froth; methods for introducing and dispersing air through nozzles, eductors, packed beds, or permeable walls; and variable cell volumes, relative to capacities, ranging from very small, with less than a second of retention time, to the largest, with retention times of over 20 min. An important question is the extent to which any of the newer designs are based on locating weaknesses in current mechanical cell designs and then using these as a starting point for innovations. In the present study, the key subprocesses which must take place in a cell are outlined; mechanical cells are analyzed to establish where weaknesses may exist; and then alternative subprocesses are considered which could effect improvements. COMPONENTS FOR CELL DESIGN It is assumed that mineral surface chemistry has been modified as necessary to provide optimum potential floatability. While the cell process may affect this to a degree, cell design is primarily concerned with actualizing the existing potential floatability. The cell design problem is to provide more effectively for the following: 1) Dispersion of air to obtain an optimum bubble size distribution at an optimum ratio of air volume to feed volume. 2) A hydraulic environment which permits most effective contact of potentially floatable particles with bubbles: [Table I - Cell Comparisons* Cell Type Mechanical Column Deister Miller ASH No. of units Multiple Single Single Single Retention time min.10/20 17/27 3.8 ca. is Capacities mt/m3s 0.022/0.066 0.066/0.20 0.11 10.4 st/ft3d 1/3 0.3//0.9 m5.0 469 Air hold-up % 15/20 20 n.a. 50/80 Air escape velocity m/s 0.015/0.0200.020/0.025 n.a. n.a. ft/m 3/4 4/5 n.a. n.a. * References: Mechanical (Harris, 1976,1985), (Bassarear, 1985); Column (Travis, et al. 1987); Deister (Zipperian and Christophersen, 1985); ASH (Air-Sparged Hydrocyclone) (Miller 1989).]

Jan 1, 1989

By R. L. Grant

[p(,ton~to'rs 'ar-IV uscG.. to explode 9crnli .... sible c}:plosivcs, dynamites, or other hiGh uxplosives. This ?,...!xplosion or detr)nation is brought about by th\.:J energy which i liberated as a cOl":"Lb' nation of concusBion and heat wl1en the ? detonator is fired. 'm Ot:18t? I'l(jrds~ liThe detonator is the key VI: ich unlocks the enerGY of a llgh? explosIve""? (8) :1/. Being sensitive nd latently dangerous, detonators ITl.ust lTIeet t\V0 primary requirerDentr: They must be ..;.afe to handle and IT. list? detonate' when (l,esired, the ... eby iIr.parting deton:l.tion to the explo8ivc cJ.1Cire-e.]

Jan 1, 1941

By B. A. Ashley

Rosebud Mining Co. (RMC) is based in Kittanning, PA. As of October 2013, it operated 24 active underground mines in Pennsylvania and Ohio. RMC?s Tracy Lynne Mine is located approximately 5 km (3 miles) east of Apollo, PA, (approximately 48 km or 30 miles northeast of Pittsburgh, PA) and mines the Lower Kittanning coal seam. The coal seam averages approximately 1.1 m (45 in.) thick and is of steam coal market quality. The mine is a single unit operation. Coal is mined using a continuous mining machine and is transported from the face to the conveyor belt using a continuous haulage system. Typical ventilation in the mine consists of an exhausting fan that draws air through isolated intake, belt and return air courses (Fig. 1).

By P. K. Naik, S. C. Das

Leaching studies of low-grade manganese ore containing 31.77% Mn and 23.71 % Fe from Nishikhal, India, were carried out using pyritiferous lignite as a reductant in sulfuric acid medium. The variables studied were quantity of lignite, sulfuric acid concentration and temperature. Regression equations for the extraction of manganese and potassium were determined from 23 full factorial design data. The experimental conditions for the extraction of manganese were optimized using the steepest-ascent method. It was found that 99.03% of the manganese could be extracted in three hours with an ore to lignite ratio of 1:1.8, a sulfuric acid concentration of 1.9 M and a temperature of 99°C. Key words: Leaching, Manganese ore, Lignite, Statistical design

Jan 1, 2003

By Yoshihiro Kawarabata

INTRODUCTION The shield-driven method in Japan has been adopted widely in soft ground tunneling, namely, subway, water supply, sewer and utility tunnels, and recently this is an indispensable method in urban area. Since this method was started to adopt on a full scale from the beginning of 1960, construction achievements in length were about 830 km for 15 years until 1974 (1). According to the recent informations of the Japan Tunneling Association, the achievement in length under construction as of December, 1977, was about 350 km (2) and the extention of increase in shield-driven method can be guessed (Tab. 1). Tab. 1 - Achievement in Length by Shield-driven Method (As of December, 1977, Unit: m) (---): Achievement by Okumura Corporation [ ]

Jan 1, 1979

By Robert F. Stevens, Robin Givet, Kenneth Viking

"Summary This paper describes a case of a directional pile installation conducted from an existing oil platform in the Caspian Sea, offshore Baku. Results are presented for 660 mm outer diameter (660 mm OD) conductor piles. Each conductor was equipped with a beveled aluminum toe for directional impact driving and was driven through a 710 mm casing using a Delmag 100–13 diesel hammer and a chaser pile. The 710 mm casings were preinstalled to a depth of 75 m below mud line (75 m BML). The longest impact driven conductor pile on the project was 456 m at end of drive, and reached the terminal depth of 295 m BML.Soil conditions consisted of soft to stiff silty clay to a depth of 30 m BML, underlain by stiff clay to 96 m BML, very stiff silty clay to 200 m BML, and very stiff to hard clay down to 300 m BML.In order to assess the maximum stresses in the conductors and check the performance of the hammer, the driving of each conductor was monitored using a Pile Driving Analyzer (PDA). The hammer fuel setting was adjusted based on the PDA outputs to ensure the conductors were not overstressed during installation. All driving and monitoring operations were carried out from the drill floor.The conductor piles drove easier than anticipated but in the direction predefined by the pre-cut angle of the beveled aluminum toe. The driving system performance was good. Post driving, the SRD were analyzed by either the global method analysis, using WEAP, or by the CAPWAP program.To detect any collision between the conductor being driven and the adjacent wells in production, the well heads were monitored in real time with high sensitivity accelerometers. These accelerometers were connected to an alarm system on the drill floor, triggering a siren and a strobe light should a collision be detected by an abnormal increase of the peak acceleration.IntroductionThe world longest directional impact driven pile was installed in year 2005. The operation was carried out from an offshore production, drilling, and quarters (PDQ) platform located about 120 km east of Baku, in the Caspian Sea.The platform foundation consists of driven piles and 24 slots were available for drilling the wells to the reservoir, located about 3800 m below sea floor. Each slot was fitted with a casing, 710 mm in diameter, running from the top of the jacket to a depth of about 75 m BML (Figure 1). Oil production started in 1997 and by the end of 2003, 14 wells were in operation; 10 slots were therefore still available, either for oil production or for water injection. However, due to larger than expected settlement of the platform, drilling of additional wells was put on hold. After further investigation, experts concluded the settlement was due to an alteration in the mechanical properties of the bearing strata supporting the foundation piles. This alteration was due to either excess soil material transportation, or soil fracture by the drilling fluids (mud), when drilling the first wells. With only 14 wells in production, out of 24 possible, the owner faced a new dilemma: to drill, or not to drill new wells?"

Jan 1, 2014

By Doug Harding

The Niagara Tunnel Project (NTP) is a 10.4-km- (6.5- mile-) long, 14.4-m (47.5-ft) bored tunnel that will run under the city of Niagara Falls from the Upper Niagara River to the Sir Adam Beck power station. The completed project will enhance the capacity of the Sir Adam Beck power station by adding 500 m3 /sec (17,657 cu ft/sec) of water through the tunnel. The project uses the world?s largest hard-rock tunnel boring machine (TBM) at 14.4 m (47.5 ft) in diameter to excavate the tunnel. Features of the project include the use of a high performance (HP) main beam TBM, state-of-the-art ground support and tunnel logistics systems. A continuous conveyor is used for muck transport from the TBM to the disposal area In June 2004, the Ontario government announced that Ontario Power Generation (OPG) had been given approval to proceed with the third tunnel under the city of Niagara Falls ? the first two tunnels were constructed in the 1950s using drill-and-blast techniques. The third tunnel, appropriately named the Niagara Tunnel Project (NTP), consists of a 10.4-km- (6.5-mile-) long TBM-bored, concrete-lined tunnel with a finished diameter of 12.5 m (41.1 ft). In August 2005, OPG the selected of Strabag of Austria, one of the largest construction companies in Europe, as the contractor to build the NTP at a price of C$600 million. In September 2005, Strabag selected The Robbins Co. of Solon, OH to design, manufacture and deliver a new 14.4- m- (47.5-ft-) diameter high performance, state-of-the-art main beam hard rock TBM to be used in the excavation of the tunnel. To meet the aggressive construction schedule as requested by OPG/Strabag, the TBM needed to be sup-plied as ready-to-bore within a 12-month period after the contract was awarded to Robbins.

Jan 1, 2008

By Frank Haehnig, Kjell-Åke Averstad, Johannes Glückert

"AbstractThe Söderströmstunneln is part of the Citybanan project that is currently under construction in the centre of Stockholm. Main part of the lot Söderströmstunneln is a 300m long immersed tunnel connecting the inner-city islands Södermalm and Riddarholmen. It consists of three tunnel elements that have been immersed in summer 2013. Since the solid bedrock is descending in that area to more then 40m depth, the tunnel elements have been placed and are permanently supported on four pile groups that have been constructed in advance. The common Swedish system of steel core piles has been used to built the pile groups. Drilling and installation of the steel cores piles in free water depth of up to 20m had to be done from a pontoon. The steel core piles are connected with underwater concrete pile caps.IntroductionCitybanan is a new 6km long double track railway tunnel running beneath central Stockholm including two new underground stations Odenplan and Stockholm City. When completed in 2017 commuter trains will have their own tracks through Stockholm, whereas the remaining railway traffic will continue to use the exisiting two tracks (see Fig. 1). This will double the capacity of the railway traffic through Stockholm. The Swedish transport administration Trafikverket is the client for this project. The design and built contract is carried out by the German contractor Züblin Spezialtiefbau GmbH together with their Swedish subsidiary Züblin Scandinavia AS.As the new Citybanan tunnel is in large parts situated below the bedrock level, it can be executed as rocktunnel in the common drill and blast method. Between the two inner-city islands Södermalm and Riddarholmen the tunnel is crossing the Söderström, an arm of the lake Mälaren. Here the bedrock is descending to a deep valley with more then 40m depth. To connect the rock tunnels between these islands a 300m long immersed concrete tunnel supported on pile foundations is currently under construction."

Jan 1, 2014