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By the courtesy of Metropolitana Milanese (Milan Underground), roost of the delegates visited on April 12, 1991 the tunnel of "PORTA VENEZIA" STATION of MILANO RAILWAY LINK. The "PORTA VENEZIA" STATION will be the most important in the link and it will he also the largest underground project of the whole railway transport network in Lomhardy. The particular conditions of the soil and the collocation of site in the center of Milano have required new technologies and the technical solution applied to this job is called "Cellular Arch Method" by its designer Prof. P. Lunardi. Our sincere thanks go to Prof. P. Lunardi and Eng. E.M. Pizzarotti of Rocksoil, Milano who have kindly furnished us a technical description of this method. Eng. Pizzarotti was the guide during the visit and gave all the technical explanations to the delegates. THE "VENEZIA" STATION IN THE 2PB SECTION OF THE RAILWAY LINK The excavation of cavities in loose ground, exceeding a certain dimension and at shallow depth, is not feasible using traditional methods: in fact it is not possible to consolidate the ground around the cavity in such a way that homogeneity and resistance is sufficiently guaranteed.

Jan 1, 1994

By Z. Q. Zhou

It is the first worke in design of pile foundation to determine the carring capacity of single pile, and a static loading on site is considered to be the lost derect and reliable way. All values P and settlements S induced can be plotted as a broken-line. Since soil is a kind of elasto-plastic and cohesive material, the increasing of settlements is generally greater than that of loads. Different value Pu nay be given by different tester on his P-S diagram, because of erroys in the test and complicated, characteristics of soil. The analytical method of exponential equation shown in this paper derives a specified solution. After using an exponential curve to regress values of loads P and corresponding settlements S from on-site test, coeficients A and B of exponential equation can be obtained. With this curve an ultimate carrying capacity of single pile can be determined. The resistant forces of pile shaft and pile tip can also be defined by the tangent of the calculated knee point, as it is called (tangent method). At last, an example is given to show how the ultimate carrying capacity of f 500mm bored pile is obtained with this method for a project in Kunming.

Jan 1, 2010

By C. H. Larsen

The results of a comprehensive investigation of conveyance arrestment devices (safety catches) in United States coal mines are presented. Availability and use, performance, inspection, maintenance, testing, procedures, and regulatory requirements were investigated. Information was obtained from equipment manufacturers, mine operators, and regulatory-agency personnel, as well as from studies of domestic and foreign mining safety regulations. Traditional mine hoists and commercial elevators for mine use were investigated. Results of this and an earlier study of wire-rope practice in United States coal mines show a continued need for conveyance arrestment devices. A wide variety of these devices are in use. Only two types have been proved reliable for stopping a conveyance without injuring passengers or equipment. These are elevator overspeed-actuated wedge-clamp (Type B) safeties and broken-rope-actuated devices for wooden guides using a single-tooth-dog design developed in Ontario as a braking element. Elevators are widely used in U. S. coal mines. When maintained according to American National Standards Institute (ANSI) Specifications A17. I and A17. 2, they appear to present no major arrestment problems. Of 42 coal-mine hoists used for man transport that were inspected, 30 are in vertical and 12 in inclined shafts. Twenty-four vertical shafts use wood guides; 6 use steel. The 12 inclined-shaft systems use flanged-wheel T-rail guidance. Two wood-guided vertical-shaft systems use the single-tooth Ontario-type dog; the other 22 use multi-toothed wedges, cams, or dogs that have been proved highly unreliable. The arrestment devices in the six steel-guided vertical-shaft systems and in six of the inclined-shaft systems cannot be evaluated because performance data are lacking. The remaining six inclined-shaft systems have no arrestment devices at all. Of the 36 hoists with arrestment devices, the devices of 16 are never tested, the devices of 10 are actuated periodically with the conveyance at rest, and the devices of the other 10 are actuated periodically with a slow-moving conveyance. Although high speed is as probable as low speed in an emergency and would provide the best test of arrestment capability, no tests are performed in the United States which demonstrate the ability of an arrestment device to stop a loaded conveyance from high speed. Mine operators indicate little confidence in their conveyance arrestment devices, and manufacturers recognize a need to improve their products. Neither state nor Federal regulations provide a guide to this improvement. Mine operators, manufacturers, and state and Federal regulatory agency personnel would welcome specific guidance. In view of this, a set of application, design, actuation, stopping-performance, and testing and inspection criteria is presented as a guide to eliminating the inadequacies of these devices. Recommendations are included for improving hoist controls and machinery directly affecting conveyance safety.

Jan 1, 1972

By K Niiranen

Luossavaara-Kiirunavaara AB (LKAB) operates one iron ore mine, three concentration plants and three pelletising plants in Kiruna, Sweden. The current methods of separation at the beneficiation plants are low intensity magnetic separation (LIMS) and reverse apatite flotation, where the wet LIMS stage is regarded as the crucial part of silica separation from the ore. Because future silica levels are expected to increase in the incoming material to the beneficiation plants, geometallurgy has become an important area of development and research.A laboratory-scale methodology was developed for the systematic characterisation of the ore for mineral processing, with the focus on the high silica ore type B2 and newly discovered subtypes. This methodology combines mineralogical, geochemical and process mineralogical characteristics. Automated mineralogy (QEMSCAN®) is used to study the modal mineralogy, the distribution of silicate minerals in different particle size classes after comminution, the deportment of silicon (Si) between various silicates and the degree of liberation and intergrowth of magnetite and silicates. An essential stage in the systematic characterisation of the ore deposit for mineral processing is the estimation and simulation of different parameters from the laboratory scale to the full scale at the concentration plants. This is achieved by using empirical models based on the data collected at the mineral processing plants. As the first step to implement the geometallurgical information, the current geological model is reconstructed with the mineralogy and the two, newly defined subtypes of high silica ore type B2 are separated into different domains.To study the reduction of the SiO2 content in the magnetite concentrate, a large-scale reverse cationic silicate flotation laboratory test program was initiated in early spring 2016. This program investigated potential silicate flotation in the beneficiation plants at the Kiirunavaara site. The second target is to implant mineral processing parameters into the block model based on the new geological model. This information provides a good basis for making a prognosis for mine planning and a quality prognosis regarding the crude ore for mineral processing.CITATION:Niiranen, K, 2017. A geometallurgical perspective on a high-grade iron ore deposit at Kiirunavaara, northern Sweden, in Proceedings Iron Ore 2017, pp 309–318 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017

By John E. Leitch

The cost of corrosion as measured by such cost components as the renovation or replacement of coated steel, or the use of expensive coating systems on steel, is considerable. Extrapolating from figures for the USA would place the overall cost of corrosion for South Africa as a figure of the order of magnitude of something over 1% of the GNP. This is a sizeable amount, of which costs associated with atmospheric corrosion would be a significant proportion. The corrosivity of an atmosphere can be defined as the ability of the atmosphere to cause corrosion in a given metal or alloy (International Standards Organization [1]). The ISO Standard 9223 specifies the key factors in the atmospheric corrosion of metals and alloys, and defines categories of corrosivity on the basis of these factors. Prior to the publication of that ISO standard, Callaghan published the results of his twenty-year exposure programme and his excellent corrosivity map of Southern Africa [2]. The present author published updates of the corrosivity map in 1999 [3] and again in 2003 [4]. The summary of the individual areas of corrosivity around Southern Africa was published in the latter paper, and is reproduced here as the following TABLE 1. Of significance in TABLE 1 are the numerous regions of very high corrosivity (C5 or beyond), particularly the rural West Coast shoreline, the North East Coast (KZN) shoreline, and the coastal cities Cape Town, Durban and Richards Bay. The corrosivity map as published is obviously based on much extrapolation as only a small number of points have been measured. During the past three years further measurements have been made, mostly however at the same sites in order to give greater reliability and to monitor long-term trends with changes in pollution. These two aspects are incorporated in this update. There has become available a greater amount of air pollution data for centres in South Africa than was the case some years ago. In this present paper the pollution data are used to extrapolate the corrosivity over areas where little or no direct corrosivity measurement has been made. This methodology is described in ISO 9223.

Jan 1, 2006

By J. E. Nesset, F. Rosenblum, J. A. Finch, R. Payant

ABSTRACT Under certain conditions of moisture and oxygen, sulphides can spontaneously self-heat. We tested a hypothesis that galvanic interaction between some sulphides can promote self-heating. Galvanic interaction is controlled by the rest potential difference between sulphides and particle size. A series of binary sulphide mixtures was tested. Individual sulphides and mixtures with low rest potential difference did not self-heat, but mixtures of high rest potential difference did. When testing pyrite-sphalerite mixtures, self-heating increased inversely with particle size, with the fineness of the pyrite governing self-heating, indicating the rate-limiting step is the reduction reaction. The results support the hypothesis that galvanic interaction contributes to sulphide self-heating. A possible mechanism based on hydrogen sulphide as an intermediate product is proposed. The understanding gained will be of interest to those involved in storage, shipping, and disposal of sulphides.

Jan 1, 2012

By P. D. Chamberlain, R. M. Nadkarni, D. J. Kinneberg

Introduction The relationship between precious metal miners and gold refiners has been uneasy over the years. Reasons for this uneasiness and lack of trust include technical factors, economics of scale, and cultural differences. Technical factors take two forms. First, concentrates, precipitates, and sponge produced at the mine site are notoriously difficult to sample. A nonrepresentative sample can be the cause of misunderstanding and distrust. Dote, however, is relatively homogeneous and can be correctly sampled and analyzed with greater reliability. Second, the technology and chemistry of fine gold and silver production is complex and quite different from dote production. Most mining companies are content not to be involved with additional processing steps. From the standpoint of process economies, it is easy to see why miners and refiners are separate groups. Mine location is dictated by the natural occurrence of ore bodies. Heavy capital investment requirements for infrastructure in remote locations usually limits companies to produce the first marketable product from a mine. Beyond that point, the economies of scale and large capital investments for refineries generally dictate that the outputs of several mines be combined to feed a single refinery. The refin¬ing process is so complex that refineries are generally built in larger cities with a skilled labor pool. Therefore, only the largest mines can afford to refine their own output. Lastly, there are cultural differences between the miner and the refiner. The miner, particularly the small gold miner, is by definition a risk taker in a risky business. Refineries, on the other hand, have historically been operated under legal or implied monopolies approved by the government. The government is interested in owning gold, either for currency reasons or because many national mints are major users of gold coins. It is not surprising, then, to find that miners may regard refiners as part of an unresponsive government bureaucracy or, worse, as "city slickers." It is easy for the two groups to distrust each other with these different backgrounds. The interface between miner and refiner does exist and affects nearly all precious metal mining companies and most refineries. This paper discusses the technical and economic aspects of the interface. It points out how both parties can improve the relationship. Technical interface Precious metals selectively dissolve to produce a pregnant solution. They are then recovered from solution by zinc precipitation (Merrill-Crowe process) or by absorption onto activated carbon. This is followed by stripping and electrowinning. At this point, the mining company must decide whether to ship precipitates (or sponge) to refiners or to produce a dote. Some factors the company must consider are the ability to sample precipitates, concentrates or sponge, security of storage, and production rates. An important factor is the incremental cost (capital and operating) of converting precipitates or sponge into dote at the mine. Generally, mines prefer to melt precipitates or sponge to dote to avoid prob¬lems associated with sampling them. There are three types of fur¬naces used by mining operators to melt precipitates to dote. The first and most widely used is a sil¬icon carbide crucible heated ex¬ternally with a gaseous liquid fuel. The crucible is easy to use and slag can be skimmed before pouring dote. This simple system is capable of handling 1 t/a (32,000 oz per year) gold. It could cost as little as $20,000. More complex systems could cost up to $200,000. The second type of furnace is the gas- or propane-fired rever¬beratory furnace. This is usually used only by large-scale opera¬tors in the silver business. Precip¬itates in reverberatory furnaces come in direct contact with the flame. The large volumes of flue gas can carry significant amounts of dust and fume that must be re¬moved from the gas before dis¬charging it to the atmosphere. In the days when the Rand re¬finery in South Africa used reverberatory furnaces, the silver leaving the furnace in the form of vapor and fume was estimated at 2.2% of the total silver charged to the furnace. The gold in the gases was about 0.03% of the gold charged. Most of this silver and gold was recovered in the exhaust gas scrubber system. These scrub¬ber systems can be more costly than the furnace itself.

Jan 10, 1986

By David R. Peterson

This report presents a deep ocean mining operational simulation which is capable of describing the response of the bottom collection unit to specified mining ship maneuvers and the corresponding operation of the mining ship maneuvering system. Bathymetry, obstacles, collector and pipe dynamics and ship maneuvering capabilities are included in the simulation. The simulation is capable of evaluating the technical requirements and operating costs for specified mine plans, Results are useful in assessing mine size requirements and provide operating 'cost 'estimates useful as inputs to deep ocean mining economic feasibility assessments, The simulation is used to determine the consequences of obstacles to mining system operations in terms of impact on mine size requirements and reduced production arising from obstacle avoidance maneuvers of a towed (unsteered) collector.

Jan 1, 1979

At the treatment plant of the Broken Hill South Limited, Broken Hill, the zinc concentrates, with water, are pumped out to a more or less level area, and stored in dams to dry by drainage and evaporation.In-the construction of these dams a clay bank is first thrown up around the proposed dam site. The concentrate and water mixture are then run in from the perimeter, the solids being continuously shovelled up to increase the bank height, while the clear water is drained off at the centre and returned to the treatment plant.These dams vary in size up to 57,000 sq. ft. in area; when completed the banks are about 12 it. high, and the capacity for the largest dams is about 350,000 cub. ft., say, 20,000 tons.The survey office is called upon for a monthly survey of these dams, generally three of which are in process of building or refilling. The puggy nature of the material prohibits walking beyond the bank of the dam, therefore the following method of estimating concentrates was evolved. No originality is claimed, but as the need is for a true survey, and the results have shown the accuracy of the method, a description is warranted. The dam site is first laid out into 10-ft. squares with a theodoliteand 150-ft. steel tape, the outside lines being placed about 15 ft. from the clay bank - 15 ft. is the maximum convenient distance from the bank for working with the levelling staff mentioned later.As the dam is rarely rectangular in plan, the outside points pegged out are not always the corner of a square, but are placed to give the greatest convenience in later work, that is to say, there may be triangles and irregular figures. The plan (Fig. 1) and photographs (Figs. 5 and 6) will give the general idea of the layout. At each point is driven a 3 ft. 6 in. round iron peg about 3 ft. into the ground. This peg is to permanently mark the point, and is not liable to be...

Jan 1, 1925

By S. B. Christy

This report on the electrodeposition of gold and silver from cyanide solutions represents work that has occupied my time at intervals during the past 20 years. The investigation has been carried on simultaneously with my duties as professor of mining and metallurgy of the University of California. During shis long investigation I have been assisted in the details of the work by, a number of my former students, acting in turn as assistants. Particular mention in this connection is due W. H. Hilton, who graduated in 1900; C. T. Dozier, 1902; N. C. Stines, 1905; H. N. Herrick, 1907, and L. C. Uren, 1911. It is impossible to publish full details of the many thousands of experiments performed. Effort has been made to present only the essential data. Wherever possible, the results of the experiments have been expressed in diagrammatic form by curves showing the relation between the simultaneous variables of the experiments. It is necessary to show, as nearly as may be, the state of the art. This is difficult because of the meager literature on electrodeposition in the cyanide process; hence resort must be had to the specifications and claims of patents, most of which seem never to have been applied on a working scale. Such information consequently can not be taken at its face value. Expressing the essential ideas evolved in such specifications is no easy task. Clearly, in quoting from patent records it is not ·feasible to enumerate all of the specifications or the claims allowed the inventors. A statement of the chief contributions to the art in the patent cited must suffice. Legal details, of course, must be obtained from the original patent. In the study of patent literature, when one compares the ardent hopes of the inventors with the results realized from their efforts, it becomes clear that in the development of an art no one man contributes everything. Each subsequent inventor stands on the shoulders of the one who has gone before. It is only through the summation of the contributions of many inventors that the art as a whole prospers. On October 19, 1887, a British patent was granted to McArthur and Forrest for the use of dilute cyanide solutions for treating gold and silver ores. The rapid extension of this process, in various forms, all over the world, has almost revolutionized the treatment of ores containing precious metals. It is a remarkable fact, however, that long before this patent was granted there was issued, on February 5, 1867, to J. H. Rae, U. S. patent 61866, for an electrocyanide process. Thus a method of electrodeposition from cyanide solutions used for treating ores was invented 20 years before the McArthur and Forrest patent .was granted, although the process did not come into general use. On May 24, 1889, U. S. patent 403202 was granted to McArthur and Forrest for the use of zinc shavings for removing gold and silver from cyanide solutions. The use of finely divided zinc is still the standard method for recovering 'gold and silver from cyanide solutions. Since this patent was granted, many attempts have been made to replace the precipitation upon the zinc by methods of electrodeposition. It is the purpose of this report to give an outline of some of these methods, as well as to show the relative advantages and disadvantages of the zinc and the electrical methods of recovering gold and silver from cyanide solutions.

Jan 1, 1919

By E. P. Youngman

This paper is one of a series of digests of foreign mining legislation and court decisions that is being prepared in advance of a general report relative to the right of American citizens to explore for minerals and to own and operate mines in various foreign countries. This interpretation of the mining laws of British Honduras has been prepared from the best available information in Washington and is released subject to amplification or correction, if necessary, by United States foreign-service officers. INTRODUCTION The mining law governing all mineral substances other than oil is ordinance No. 7 of 1927 (Minerals Ordinance, 1927), passed by the Legislative Council on March 25, 1927. The accompanying regulations, No. 3 of 1932 (General Minerals Regulations, 1932), made under the authority of section 66 of the minerals ordinance, were made by the Governor in Council on December 23, 1931.

Jan 1, 1934

By L. J. E. Hofer

THIS BULLETIN analyzes and summarizes research at the Bureau of Mines dealing with the preparation, properties, and reactions of iron carbides in the period 1948-60, part of which was sponsored by Wright Air Development Laboratories. The objective was to develop better catalysts for the synthesis of methane and higher hydrocarbons from hydrogen and carbon monoxide. The investigations resulted in the first adequate characterization of epsilon iron carbide, which later proved to be an important intermediate in the tempering of martensite in steel. At temperatures near 170° C, the carbon in carbon monoxide reacts nearly quantitatively with finely divided alpha iron to form an iron carbide in which the metal atoms occupy the positions of the close packed hexagonal arrangement with lattice parameters of ao=2.754 A and co=4.349 A. The carbide has a composition within 2 atomic percent of that required for the formula Fe2C, a Curie point of about 370° C, a specific magnetization of 160.4 ergs per gauss-gram Fe at 50° C, and a ferromagnetic moment of 1.70 Bohr magnetons. Epsilon iron carbide can be converted to chi iron carbide by thermal treatment at 343° C for about 70 hours. The chi iron carbide so formed has a Curie point of 247° C, a specific magnetization of 142.5 ergs per gauss-gram Fe at 50° C, a ferromagnetic moment of 1.72 Bohr magnetons per iron atom, and a paramagnetic moment of 5.5 Bohr rnagnetons per iron atom. Direct synthesis of this carbide from carbon monoxide at 240° C indicates a composition within 7 atomic percent of that required for Fe2C. Chi iron carbide can be converted to cementite at 550° C in about 315 minutes. The theta iron carbide so prepared has a Curie point of 210° C, a specific magnetization of 147.3 ergs per gauss gram Fe at 50° C, a ferromagnetic moment of 1.72 Bohr magnetons per iron atom, and a paramagnetic moment of 3.9 Bohr magnetons per iron atom. Pure chi iron carbide reacts to form theta iron carbide by the elimination of free carbon. In the presence of an excess of free iron in intimate contact with the carbide, chi iron carbide reacts with the free iron to form theta iron carbide. The second reaction proceeds at an appreciable rate at 320° C and can be used to snythesize finely divided theta iron carbide with no free-carbon impurity at low temperatures. The ease of the reaction in the presence of free iron suggests that chi iron carbide is an intermediate in the tempering of epsilon iron carbide to theta iron carbide as it proceeds in steels. Experiments on carbon steel show that the magnetic moment of the iron atoms in martensite is about 2.2 Bohr magnetons per atom, a value approximately the same as that of the iron atoms in alpha iron. The Curie points of epsilon iron carbide, chi iron carbide, and theta iron carbide were found -in various specimens of tempered steel to be 10° to 20° C lower than normal values. This shift is probably the result of manganese in solid solution. Preparations of various compositions of epsilon iron carbonitrides show many effects of extensive isomorphism between epsilon iron carbide and epsilon iron nitride. The action of carbon monoxide on finely divided alpha iron at 400° C results in free carbon deposits containing filaments about 0.1 micron in diameter and several times as long. The filaments are characterized by nuclei of dense material located midway from each end of the filament. Although carbiding to near completion at 240° C over periods of about 300 hours produces a preparation containing no appreciable percentage of epsilon iron carbide, very brief periods of carbiding of the order of a minute result in the formation of detectable epsilon iron carbide even at temperatures as high as 325° C. The data indicate that epsilon iron carbide is an intermediate in the formation of chi iron carbide during its formation from carbon monoxide in the gas phase. Attempts to convert cementite to chi iron carbide by cold work increased the Curie point but did not change the X-ray powder diffraction pattern.

Jan 1, 1966

On June 20, 1972, Ocean Energy was awarded Contract No. H0220055, entitled Portable Task Luminaire System Development. Initially, this was a 12 month program consisting of three phases: Phase I -- Design of the Luminaire System and Battery Chargers, Phase II -- Fabrication of Prototype Units, Phase III -- Fabrication of Production Prototype Units and Battery Chargers. In addition to developing the Portable Task Luminaire System, however, this contract was amended a number of times in the course of the program to include the development of a Low Coal Portable Task Luminaire, a Rib Mounted Luminaire System, and a Machine Mounted Luminaire System, all based upon the same high frequency fluorescent lighting and safety technologies involved in the Portable Task Luminaire. This report covers the activities of this program as amended, and includes the development and fabrication of all of the systems. This report constitutes the final fulfillment of the contract.

Jan 1, 1974

By Paul J. Lewis

Estimation of the axial side shear resistance of drilled shaft rock sockets is directly related to the unconfined compressive strength of the bedrock. Commonly, the mean unconfined compressive strength value is calculated from a range of laboratory test results and is used to estimate the side shear of the bedrock. This paper presents the results from a full scale axial load test performed on an 8.5-feet-diameter drilled shaft rock socket, and compares measured side shear values to estimated values for two different rock strata. Measured side shear values correlated well with the design value calculated for the "strong" rock stratum, but did not correlate well with the design value calculated for the "weak" rock stratum. The mean unconfined compressive strength was used to calculate the side shear in both strata, but review of the laboratory data indicates that this value may not have been appropriate for the weak stratum due to a large variation in the laboratory test results. Also included in this paper is a discussion of the subsurface conditions, the laboratory test results, and the methodology used to design the drilled shaft rock sockets for axial and lateral resistance.

Jan 1, 2005

By Paul S. Goodman

This report examines the feasibility of Employee Assistance Plans for the coal industry. Three issues are examined in detail. First, the prevalence of off-the-job factors such as alcoholism, drug abuse, and marital problems are examined in a sample of miners from the eastern, midwestern, and western underground coal fields. Then data were presented that examined the link between these off-the-job factors and behaviors critical to the mines such as absenteeism and turnover. Lastly, we examined the effectiveness of the major Employee Assistance Program in underground coal mining--the HELP Program. The report contains recommendations on whether coal mining companies should implement these programs and how these programs should be designed.

Jan 1, 1986

By Peter Radziszewski

"Comminition process costs, which typically represent an important contribution to mining operating costs, can themselves be divided roughly between energy and liner /charge wear. As such, developing the ability of predicting mill wear will allow a more accurate estimation of wear costs and therefore their control. In this paper and based on the possibility of estimating energy dissipated in impact and abrasion by using a discrete element charge motion model for mill charge motion simulation, media wear will be investigated principally as a function of three wear mechanisms : impact, abrasion and corrosion. Experimental results will be used to identify a total media wear model. Model results will be compared with real mill data from three operating Australian mines. These preliminary results show for the cases tested that the overall average error, as compared with the observed cases, is 0 % with a standard deviation of 66.1 %. Results from the Bond abrasion index tests, for the same cases investigated, give an overall error of 73 % and a standard deviation of 192.5 %. Further work is in progress to improve test procedure, validity and confidence.INTRODUCTIONComminition process costs, which typically represent an important contribution to mining operating costs, can themselves be divided roughly between energy and line / charge wear. In Greenfield situations where the estimation of operating costs is critical in determining project feasibility, the level of confidence in such estimates becomes quite important. With regards to media wear, the Bond abrasion test currently serves as the standard media wear prediction device. However, as shown [l, 21 test results once compared with actual wear rates can vary greatly. Figure 1 shows test results as compared with 15 cases. Average error over this data is 73 % with a standard deviation of 192.5 %. As such, developing the ability of more accurately predicting mill wear will allow greater confidence in wear cost estimation and therefore their control."

Jan 1, 2001

By Jerry J. Ran

In open stope mining borehole deviation is one of the main factors affecting the blasting results. Though previous research has identified some actual deviation patterns, these studies are unable to relate the hole deviation pattern with the blasting results and therefore it is not possible to modify the design or take remedial measures based on the prediction of the blasting results. An alternative to the conventional method for borehole surveying is the in-hole probing techniques. With their inherent errors, it is important to evaluate these surveying systems under conditions of different mine sites. The lack of available assessment methods has limited the application of these techniques. In this paper some of the characteristics of the actual hole deviation pattern are discussed based on data from surveying production holes in six stopes. A method of predicting the blasting results is proposed by considering the borehole deviation pattern. Based on the prediction stope design can be modified or further surveying is carried out to improve the blasting results. The tolerable error is introduced to relate the surveying error with the drilling pattern and therefore it is possible to assess various in-hole-surveying techniques.

May 1, 2002

By L. Maday, V. J. Perrone, M. B. Gilbert

INTRODUCTION The Brightwater conveyance system project in King County, Washington consists of constructing about 21 km (13 mile) of effluent and combined influent and effluent tunnels. The influent tunnels will collect sanitary sewerage from the service area and convey it to the new wastewater treatment plant (WTP) site at Route 9 in Woodinville, Snohomish County, Washington. The effluent tunnels will convey treated effluent from the WTP to an outfall at Puget Sound. The tunnels will be constructed primarily under residential areas, busy suburban roads and state highways. This 21 Km (13-mile) soft ground tunneling project faced many design challenges because of the geologic and hydrostatic conditions, long tunnel reaches, and large overburden depths. Boring depths typically varied from 45.7 M (150 ft) to 114 M(375 ft) with a maximum depth of 185.6 M (566 ft) deep. Conventional empirical methods for non-mountain terrain based on a 1:1 ratio of drilled footage per foot of tunnel bore (National Research Council, 1984) for evenly spaced borings provided a good start for the Phase 1 explorations. However a continuation of this approach with closer spacing was deemed inadequate by the geotechnical team for efficiently allocating geotechnical exploration resources for the final design. This paper presents a rational approach that was used to evaluate the value of a particular geotechnical exploration boring, pump test or other in-situ test. The method of analysis, which was initially developed for rock tunnels (Einstein et al., 1978), was modified for the Brightwater project. The method considers the potential construction cost risks resulting from variable geologic conditions, the cost and reliability of the geotechnical exploration and the earned value (cost risk reduction) of performing the geotechnical exploration. PROJECT DESCRIPTION In total the land portion of the conveyance system will include approximately21 Km (13 miles) of tunnels ranging in size from 3.0 to 5.2 M (10 to 17 ft) minimum inside diameter with overburden depths of 30.0 to 168 M (100 to 550 ft) and

Jan 1, 2005

By P. F. Allen

WITH capacity of 51,000 tpd, the Morenci concentrator produces approximately 49,000 tons of tailing for final deposition. Disposal involves distribution of thickened tailing to dams in such a manner that solids can be used for dam construction and water can be impounded for reclamation. During early operation, although tonnage treated was lower, it was necessary to use orthodox disposal practices in the areas available for tailing. Trestles mounted on a series of dams, constructed of borrow material, carried the pipelines from which tailing were spigoted behind the dams. When adequate disposal areas were finally established, it became desirable to look for more efficient methods. The idea of placing the distribution pipe or manifold directly on the crest of the dam was advanced in 1945. After a period of testing, that procedure in modified form replaced the older trestle arrangement. Marked economies resulted, and time required to prepare dams for succeeding storage cycles was greatly reduced. The tailing disposal area for the Morenci concentrator lies south and southwest of the concentrator building. Ample storage space is available. A wide valley contains three sharply defined ravines lying on an average slope of 4 pct. The more central of these ravines, and the largest of the three, is known as Morenci canyon, which provides substantially more tailing storage area than the other two. Stargo canyon lies to the west; the other is appropriately named East canyon. Total area of the tailing disposal site is nearly 31/4 sq miles, or about 2000 acres. The storage area in use was somewhat more than 850 acres at the end of 1954, or 16 acres per 1000 tons of tailing to be stored daily. At the present time none of the east side dams are in regular use but they are available in emergency. There are five dams on the west side (in Morenci and Stargo canyons) as well as two auxiliary dams adjacent to No. 4 West and No. 5 West. This has made it possible to concentrate men and equipment in a smaller area with consequent savings in both manpower and machinery. Tailing Thickening: Tailing from the flotation process reaches the thickeners with a solid content of about 18 pct. The launder carrying this pulp is subdivided at a central point into seven launders carrying a proportional share to each of seven tailing thickeners, which are located in Morenci canyon close to the concentrator. Five of these thickeners are 300 ft diam and two are 325 ft diam. Thickeners are of the peripheral traction type and rakes travel at the rate of one revolution in 37 to 39 min. The entire installation provides 500,000 sq ft of settling area, or approximately 101/2 sq ft per ton of tailing to be thickened daily.

Jan 1, 1957

By Edwin W. Meyer

The augered cast-in-place piling technique was used to install deep foundations 20 feet into weathered sandstone. The project involved the upgrading of the Air Quality Control System (AQCS) for a 560 megawatt coal fired power plant located in Stillwater, Minnesota. The site subsurface consisted of 20 feet of overburden overlying weak sandstone. The overburden consisted of sands and gravels with cobbles and boulders, and variable thicknesses of peat. Ground-water at the site was approximately 8 feet deep. Two foot diameter piers were selected to facilitate installation through the cobbles and boulders in the over-burden and to penetrate the sandstone 22 feet. During installation of the first pier, the weak sandstone broke down to sand-sized material that could not be cleaned from the pier hole without the use of slurry to hold the material in suspension. However, the use of slurry would significantly impact both schedule and cost. Since the sandstone could be easily drilled with the pier equipment, it was believed that 2 foot diameter deep foundations could be installed as designed using augered cast-in-place piling techniques. A brief test drilling pro-gram was conducted to evaluate the viability of 2 foot diameter continuous flight augers penetrating 22 feet into the weak sandstone. The test drilling indicated that augered cast-in-place installation techniques should be successful. Following a short delay to mobilize the additional required equipment, construction commenced. Installation of the 2 foot diameter augered cast-in-place piling to a depth of 22 feet into the weak sandstone was successfully performed.

Jan 1, 2006