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By Yifan Ding, Shemek Oginski, Patrick Finn

The Manhattan section of the New York Water Tunnel No. 3 constructed by the Schiavone/Shea/Frontier-Kemper Joint Venture is a $668 million project. The contract includes a tunnel drive of 8.85 km (29,045 feet), tunnel lining of 14.77 km (48,468 feet)and construction of nine shafts ranging in depth from 143 m (469 feet) to 169 m(555 feet). Construction methods for excavation of the shafts are presented including raise bore, ground freezing, soil excavation, blasting, slashing, muck handling, rock support and shaft final concrete lining.

Jan 1, 2007

This report is the 25th in a series summarizing Bureau of Mines research and technologic work on coal and related investigations, and the 7th of these reports to be issued on a calendar-year basis.3 It reviews briefly work in progress during 1960, and furnishes references to publications by Bureau personnel during 1960 and also to a few older papers that were not cited in prior reports of this series. Details about some of the researches mentioned can be found by referring to the publications cited in the footnotes or in the further list in the Miscellaneous section; other phases of work are reported here for the first time and will be covered in more detail by later publications, Coal programs of the Bureau of Mines extend over the wide range of theoretical and applied studies indicated by the table of contents. Fundamental research on coal structure and on the properties of coal will, over a period of time, increase the general knowledge about coal Which scientists and engineers may use in answering practical questions and in solving problems in the mining and utilization of coal. More applied studies are conducted in connection with the developement and evaluation of new mining techniques and the effective processing and utilization of coal and of coal products. Health and safety activities, analyses of coal being mined, and technical advice and assistance to Federal agencies in connection with fuel and boiler-water problems are among the practical services handled by the Bureau.4 5

Jan 1, 1962

By F. Calizaya

The Western Mining Training and Translation Center is a consortium of four universities devoted to developing educational and training projects to reduce the incidence/accident rates in the mining industry. The center consists of Missouri University of Science and Technology, Montana Tech, the Colorado School of Mines and the University of Utah. It is fully funded by the National Institute for Occupational Safety and Health (NIOSH). The center manages 15 educational projects and one translation project. Hazard Identification and Risk Assessment (HIRA) for Small Mines is an educational project managed by the University of Utah. The objectives of this project are to identify critical mine hazards, assess risks; outline procedures to reduce accident/incident rates in small mines in the western United States and to train mine workers and super-visors on HIRA in each participating mine. To accomplish these objectives, the HIRA personnel visited 15 small mines in Utah, Colorado, Nevada and California. At these mines, most operators were employing independent contractors to perform several tasks ranging from drilling, blasting and mucking of rock at the work site to hauling ore to a processing plant. The operation of diesel-powered equipment was one of the major tasks performed by the contractors. This article presents three case studies in which contract workers performed a number of tasks with no reportable incidents but did so unsafely. It also highlights the training received by most contractors, the constraints imposed upon them and the safety hazards to which they are exposed most frequently.

Jan 1, 2008

By J. N. Van der Merwe

In the process of pillar extraction, pillars are seldom if ever extracted completely. The pillar remnants, or snooks, play an important role in the extraction process. At the working face, they need to be stable to provide a measure of support in the immediate area. Further back, they have to fail in a controlled manner to prevent undue load build-up on the pillars that have not yet been extracted. The stability condition at the working area is the result of interaction between the stability of the pillar and that of the roof rock. It is shown that the primary mode of roof layer failure is by bending induced tension. Direct tension can be excluded in most practical conditions, and only in exceptional cases will the first failure be in shear mode. Panel spans required to cause the first goaf to be formed, predicted using analytical considerations, are of the same order as has been observed in practice. Once the first goaf has formed, the horizontal stress is reduced and failure will occur at more frequent intervals. It is shown that bending induced tension is the most likely mode of failure in an unjointed roof layer. In roof layers with a high frequency of weak or uncemented jointing, failure can be expected to occur at spans equal to the joint spacing. In this context, failure will occur at the smallest of the spans required to induce tensile roof failure and the joint spacing. A fundamental procedure to evaluate the likely roof layer failure in a stacked beam system is developed. This can to be used to estimate the loading condition on snooks. It is shown that snooks assist in stabilizing the immediate roof beam. The sizes and positions of the snooks relative to the position of the next pillar to be stooped, control the magnitude of the contribution to stability. The further the snook is from the next pillar, and the larger it is, the greater the contribution to the stability of the immediate roof. Under the condition of loading by the immediate, continuous roof beam, the position and size of a single snook controls the load acting on it. The closer it is to the nearest solid, provided that the roof overhang extends beyond the snook, the greater the load on it. However, once the immediate roof beam fails, the loading conditions change. The snook is then loaded by the tributary area load. A procedure is developed to assess the condition of snooks for different situations. Using the fundamental considerations developed in this paper, snooks can be designed to be stable during the active mining phase in their immediate vicinity, but to fail once the next line of pillars has been extracted.

Jan 1, 2005

By R. H. Seraphim

"Western Mines Limited produces gold, silver, copper, lead and zinc from a number of Paleozoic volcanogenic deposits near Buttle Lake on Vancouver Island. The deposits were emplaced along a series of submarine linear volcanic ridges that extended for at least 6,000 metres in a northwesterly direction.Rock types mapped include: andesitic, dacitic and rhyolitic flow s and volcanic breccias; detrital breccias, tuffs and hematitic silts; black and green cherts; and a variety of dykes and sills, including quartz porphyry and diorite. Orebodies are associated invariably, and in places intimately, with quartzsericite schist and rhyolitic rocks. The ore is complex, varying in mineralogy and tenor both within each deposit and from one deposit to another.Recent studies have increased our knowledge of the structure and left no doubt that the orebodies were deposited both within the throats of the linear vents and on paleoslopes adjacent to these vents.IntroductionWestern Mines Limited operates an 850-tonne-per-day mine, located 85 km south of Campbell River, B.C. The mine is at approximately 400 metres elevation, near the south end of Buttle Lake on Vancouver Island (Fig . I). It has operated since early 1967 and produced 3,650,000 tonnes with an average grade of 1.9 g (0.06 oz) gold, 93 g (3.0 oz) silver, 1.6 per cent copper, 1.0 per cent lead and 7.5 per cent zinc per tonne by the end of 1978. Current reserves are 1,157,000 tonnes of 2.7 g (0.09 oz) gold, 117 g (3 .8 oz) silver , 1.2 per cent copper, 1.2 per cent lead and 8.0 percent zinc per tonne (December 31, 1978).Ore from one zone, known as Myra High-Grade, is mined and milled separately because of its grade. Ore milled prior to December 31,1978, plus reserves, 180,715 tonnes of6.6 g (0.21 oz) gold, 554.5 g (17 .7 oz) silver, 0 .8 per cent copper, 3.2 per cent lead and 10.6 per cent zinc are included in the figures listed in the preceding paragraphs."

Jan 1, 1980

By R. L. Amelunxen

"In January 1983, Gibraltar Mines Ltd. started an exploratory program on Column Flotation with hopes of improving performance of the Molybdenum Cleaning circuit. The encouraging results led to a column test trial in the Bulk Copper Cleaner circuit where the test unit has per¬formed well. A program was undertaken to understand the mechanics of column operation. A year and a half later, Gibraltar is considering the replacement of existing Bulk Copper and Moly Cleaning flotation cells with a fully computer con¬trolled Column Flotation circuit.BRIEF SUMMARY OF TESTWORKInitial testwork was done with a home made 20 inch diameter x 31 foot high column, equipped with a doppler flow-meter at the tail line and automatic level conrol. This loop maintained constant level in the column by throttling the tail valve. This column was placed in a variety of locations around the Moly circuit as shown on Figure 1 a, b. Test re¬sults indicated that the column could achieve much higher grades than the conventional cells, but recovery wise, the performance was low. Spillages from overloaded pump boxes were a common frustration during this time.Column control strategy was reviewed; and, in light of information received from McGill University, changed to conform with the methods that had been employed by D. Wheeler, of the Column Co. of Canada4 and Gaspe Copper Mines3.The column was tested replacing a portion of the bank, figure lc. The results obtained were more or less equal to the circuit. However, as we had Nitrogen gas connected to the column and not the regular cells, the column was kept in service. With the use of Nitrogen in the column, the usual copper depressant, Sodium Cyanide could be replaced with Sodium Hydrosulphide, hence lower operating cost.As time went on, a 10 inch diameter column was added in series to re-treat the concentrate of the 20 inch diameter column with hopes that the copper content would decrease from 0.6% down to 0.3% in order to save on subsequent leach¬ing costs. So far, results indicate there has been some im-provement, hut we are not achieving 0.3% copper. (See figure 1d)."

Jan 1, 1985

Three stages are involved in the complete design and evaluation of an open cut mine. These are:1. design of the optimu,m ultimate pit, and selectionof the optimum mill capacity,2. development of production schedules, with intermediate pit designs, and3. evaluation of the project, based on factors from stages 1 and 2, by means of discounted cash flow analysis.Stage 1 requires relatively complete information on metal distribution, metallurgical performance, and cost factors, which is used to design a series of pits, corresponding to a range of milling rates.For each milling rate considered there is a unique pit design which maximizes operating profitability. The computer outlines this design (in terms of size, shape, and location; tons and grade of ore; and tons of waste), by considering the effect on profitability of mining each block in the mineralized inventory in relation to every other block which would be affected by its removal.The computer selects only those blocks for mining which result in an overall increase in profitability, maintaining pre-determined wall slope conditions throughout.Both operating profitabilities and capital costs increase with throughput, and the two curves will combine to produce a flat peak in net profitabilities at a certain rate, with its associated pit design.Stage 2 can maximize head grades early in the life of the operation. This increases the projected present value, but there are good arguments for maintaining steady head grades throughout the life of the mine.A modification of the programme used in Stage 1 is used here.Stage 3 is a conventional discounted cash flow analysis, using factors obtained in stages 1 and 2. It can be computerised, but a manual treatment is probably just as fast overall.

Jan 1, 1970

By W. A. Scheuch

ORDINARILY secondary copper, unless electrolytically refined, is reclaimed directly as foundry ingots used in the manufacture of copper-rich alloy castings. This use does not require the elimination of the impurities to the degree required if the recovered metal is to be used for electrical purposes where a high electrical conductivity is of prime importance. However, as a part of the general policy of reusing its own scrap and by-products wherever possible, the Western Electric Co. at its Hawthorne Works is recovering its secondary copper as wire bars which are delivered to its own rod and wire mill and there processed into wire of high electrical conductivity. This departure from the usual recovery scheme has been accomplished through the careful classification and sorting of the material entering the furnace charges and the careful control of the subsequent melting and refining in a reverberatory furnace. Tests conducted over a period of approximately four years have shown that the quality of wire bars produced from selected secondary copper is equal to that of wire bars produced by the electrolytic refineries. The interest which has been exhibited in the process by those who have seen it has prompted the following general description. It is the authors' opinion that the recovery of secondary copper will become more and more an important factor in the copper production industry. The description is largely of the refining furnace and its operation, but it must be remembered that careful classification and sorting of the scrap materials to eliminate impurities are just as important as the furnace operation toward obtaining a satisfactory final product. Classification and sorting do not lend themselves easily to description; they must just be done. It should also be recognized at the outset that constant intelligent supervision is requisite to the success of the process because there are so many ways in which impurities can find their way into the final product.

Jan 1, 1929

1. Proposals were submitted to the Western Administrative Center of the US Bureau of Mines, Denver, Colorado on 30th December, 1974 and Contract No J 0255003 was awarded, the effective date being 8th May, 1975. Dames and Moore, Denver, are associated with PD-NCB Consultants in regard to environ-mental problems, surface mining practice and legislation in US. BACKGROUND TO THE STUDY 2. Under the threefold pressures of -increased domestic fossil fuel demand -increased value of domestic energy resources -increased demand for environmental protection and safety

Jan 1, 1976

By D. P. Venter

When Oryx Mine down-scaled in 1994, focus was placed on development, with the primary aim being the exposure and proving of ore reserves. In order to achieve this aim, development ends had to be advanced to reef at the maximum possible rate. This in turn called for support methods which could keep up with the advancing faces, taking up minimum space and time for installation, while providing on the face a real coverage and effective long-term support. In the majority of the development ends which were located in fairly competent ground, the grouted rod development support proved to be adequate until such time as secondary support in the form of meshing and lacing could be installed. The ends developing towards the eastern boundary of the mine, especially in the ?C? block, encountered extremely adverse ground conditions due to the presence of smectite zones within the footwall quartzites. Smectites are a group of montmorillonite clay materials, mainly composed of hydrous aluminosilicates. They occur as minute platy crystals in the rock mass and are especially notable for their ability to absorb liquids. Generally, the zones are found in between the bedding planes of the host rock, and may vary considerably in thickness over short distances. Wherever the smectites were exposed by the development blasting, the rockwalls weathered as the smectites absorbed water and the condition of the tunnel rockwalls deteriorated rapidly. When the smectites were left exposed and allowed to weather, it became almost impossible to install secondary permanent support such as meshing and lacing at a later stage. To minimise the exposure time and subsequent deterioration of the smectite zones, it was recommended that all development ends encountering smectites be shotcreted as soon as possible after blasting. In developing a shotcreting support methodology, both a drycreting and wetcreting system were investigated. The situation at the time ruled out the use of a wetcreting system on the basis of the extensive and costly infrastructure required, as well as its limited manoeuverability. Instead, it was decided to opt for as maller scale but more manoeuverable drycreting system. The system utilized a number of portable drycrete machines with pre-bagged cement, which could support several development ends in close proximity. Initial trials using normal 30 MPa shotcrete on 19level proved successful and the system was implemented.

Jan 1, 1998

In 1801, Hatchett discovered the oxide of a new element while he was examining a heavy black mineral from Haddam, America, which had been presented to the British Museum. He named the mineral Columbite and the metal Columbium. A year later a Swedish investigator, Eckeberg, while working with a new mineral resembling columbite-yttrotantalite discovered a new element, to which he gave the name of tantalum.In 1844, Rose announced the existence of two new elements in a sample of columbite from Bodenmais. One was similar to Eckeberg's tantalum and the other was called Niobium. Wollaston, in 1909, attempted to show experimentally that columbium and the tantalum were identical, but it was soon proved that they were two separate elements. Subsequently a number of investigatiors found other "new elements" in tantalum and columbium bearing ores, but these were shown to be mixtures of the twin metals.Since Hatchett's Columbium and Rose's Niobium are identical, it is more fitting to use the name assigned to this element by the earlier investigator. However, the term niobium is still retained in Germany and England.The separation of tantalum and columbium has been one of the major problems for analytical chemists since the discovery of the two elements and it is only in recent years that tantalum of a high degree of purity has been produced. High purity columbium metal is probably still to be obtained.Until 1903 tantalum was merely a laboratory curiosity, but about this time its use in incandescent electric light bulb filaments was discovered. Eight years later it was replaced by tungsten, which was better and cheaper. For a time the electrolytic valve action of tantalum was used in the charging of storage batteries, but the chief uses of the metals' to-day are for corrosion resistant apparatus, in vacuum tubes and as carbides in the manufacture of hard cutting and shaping tools.Most of the recent work on the metals has been done by C. W. Balke, in America. Tantalum has been studied more than columbium. OCCURRENCE OF TANTALUM AND COLUMBIUMI. NATURE OF ORESTantalum and columbium occur, nearly always together, as the oxides in combination with other metallic oxides. The only known exception to this is the reported occurrence of metallic tantalum with gold in the Ural Mountains. Unlike most other metals, no compounds with sulphur are known to occur in nature. Tantalum and columbium minerals occur usually...

Jan 1, 1943

By K C. Williams, A W. Roberts, J D. Prigge

For stockpile design, the flowability and structural interfaces are critically dependant on the stress conditions throughout the pile. Historically, the predictive methodologies for stress determination have been too conservative, which adds inefficiency to the stockpile reclaim interface selections, especially for live reclaim stockpile systems. In addition, accurate pressure profiling throughout stockpile is essential to correctly design the stockpile support structures, which are buried within the stockpile. Such buried structures include trestle legs or columns to support load-out conveyors or the roof of bulk storage sheds. Improved methodologies with respect to the development of the hoop stress condition which exists in material ædrawn-downÆ within the stockpile has been developed. This hoop stress approach has significantly improved the live capacity prediction to over 90 per cent accuracy and improved rathole structure prediction within live reclaim stockpiles. Additionally the methodologies in estimating the vertical and circumferential stress variation acting on buried stockpile structures have been significantly progressed, with experimental testing shown good agreement with the theory.CITATION:Williams, K C, Roberts, A W and Prigge, J D, 2013. Recent developments in stress analysis and live reclaim prediction within iron ore stockpiles, in Proceedings Iron Ore 2013 , pp 501-510 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 12, 2013

By G. Richards

Research shows that having more women in the top mining jobs can bring bottom-line benefits. Mining companies are starting to embrace the need for greater gender diversity at board and senior management levels, figures and examples are given.

Jul 1, 2015

By Jan E. Nesset, Colin Hardie, Peter Radziszewski, Donald P. Leroux

"The successful introduction of new high intensity stirred milling technologies (e.g. IsaMill, SMDetritor) to the mining industry in recent years prompted Noranda-Falconbridge to initiate a study comparing these technologies to the conventional technologies of ball milling and existing vertical stirred mills (e.g. VertiMill). The study was unique in that it utilized the same material (Brunswick Mine zinc cleaner regrind mill feed), the same sizing instrument (a Malvern laser difractometer) and sampling protocol, and media size and type (for the high intensity stirred mills). The paper compares the net energy efficiency (kWh/t) of the various technologies, their power intensity (kW/m3) and product size distribution spread (P80/P20). A key finding is that media size is critical to achieving optimal performance. As part of the performance comparison, an analysis of the four methods of power measurement was made (Prony brake, shaft torque, grinding chamber reaction torque, “no-load” correction) and followed-up with an experimental test procedure using three of these methods. The findings include divergent results for power measurements and some explanations based on the observations are proposed. The direct shaft torque and Prony brake calibration methods are shown to be the methods most suitable for stirred mills.INTRODUCTIONThe motivation for the investigation reported here was the lack of a suitable comparison within the available literature of relative grinding efficiency for the emerging crop of new industrial fine grinding technologies. Such comparisons are potentially fraught with inconsistency, from ore samples and grinding media used, to testing methodology, to size measurement techniques, to the power measurements themselves. The collaboration of the industrial suppliers (at the time MIM Technologies and Svedala, now Xstrada Technologies and Metso Minerals) and McGill University (P. Radsiszewski) were sought in order to ensure as valid a comparison as possible. The work was performed between October 2000 and November 2002, led by the mineral processing group at the Noranda Technology Centre (NTC) in Pointe-Claire, Quebec. The paper is presented in two sections; Part A considers the different power measurement methods and a comparison of the results for the SMDetritor (SMD) technology, while Part B reviews the grinding performance of the 5 technologies investigated using the selected performance criteria."

Jan 1, 2006

By T. Tamagawa, N. S. Fu, I. Iwasaki, M. Kobayashi

A 3-stage fluidizing bed process of chlorination, selective oxidation and sulfur chloride conversion was tested for processing of a chalcopyrite concentrate. It was found that the use of excess oxygen was essential to a stable operation in the oxidation reactor. The oxygen-excess gas will inevitably cause the oxidation of the sulfide concentrate in the chlorination and sulfur chloride conversion reactors. A new flowsheet is proposed to separate the chlorination and the selective oxidation steps with separate flow paths of the fluidizing gases, thereby preventing the passage of oxygen-excess gas in the chlorination and sulfur chloride conversion reactors. The feasibility of the proposed flowsheet was demonstrated in a laboratory-scale fluidizing bed setup.

Jan 1, 2000

By C. Bazin, D. Hodouin

"A flexible model involving perfect mixers, stagnant zones, plug flow, material bypassing and recycling zones has been developed for interpreting mixing phenomena in mineral processes. The mixing properties are described by the residence time distributions which can be subsequently interpreted by the flexible model. User-friendly computer programs (for 1MB-PC or compatible) have been developed for simulation and modelling of RTDs using the proposed model. Seven mineral processing case studies are discussed in terms of experimental procedure, model calibration results and application of the estimated mixing model to process modelling, diagnosis, and performance evaluation. IntroductionIn any mineral processing plant there are continuous-flow machines that transform or separate the ore components. Grinding mills, flotation cells, conditioning and leaching tanks are examples of such machines. The performances of these vessels or reactors are controlled by the rate at which the reaction proceeds and the time spent in the environment where the reaction occurs. This paper deals with the latter aspect of mineral processes operation.Since most of the reactions occurring in mineral processing units are time dependent, knowledge of the residence time of the material in process vessels is important for modelling and diagnosis purposes. Due to the random mixing phenomena involved in the process equipment, there is a distribution of the residence time of the ore particles and the fluid in the vessel. The probability density function E(t) that a micro-element of the material flowing through a processing unit stays a time t in the vessel, is the residence time distribution (RTD) of the corresponding component."

Jan 1, 1988

By J. R. Ross

A reconnaissance for sources of scandium and a metallurgical investigation of recovery methods for byproduct scandium from uranium mill solutions were conducted by the Bureau of Mines. The work on scandium was undertaken to obtain information on domestic resources and to explore and develop recovery procedures in consonance with the Bureau's long-range program for conservation and wise utilization of mineral resources. The element scandium is widely distributed in the earth's crust in trace amounts, but it seldom occurs as an important constituent in any mineral. An exception is thortveitite, (Sc,Y)2 Si207) a rare mineral that contains various amounts of scandium and yttrium and may contain up to 42 percent scandium oxide (SC203). Knowledge of the chemistry of scandium has expanded rapidly in recent years (1, 4, 9)4 but only fragmentary information is available about the metal because it is still largely a laboratory curiosity without established uses (2, 3). A recent study shows some of the physical proper-ties of the metal are (1) density of 3.05 grams per cubic centimeter) (2) melting point of 1,540° C, (3) hexagonal close-packed structure) (4) hardness of about 50 Brinell (500-kilogram load), and (5) an alloy behavior similar to yttrium and rare-earth metals (2). The price of $2,850 a pound for scandium oxide of 99.9 percent purity and $2,700 a pound for oxide of 99 percent purity reflects the scarcity and complex chemistry of the element.

Jan 1, 1962

By Peter J. Nicholson

Since the reintroduction of soil mixing to the United States over 10 years ago, the work installed by soil mixing technology has grown dramatically. From an annual turnover of less than ten million dollars per year the volume of soil mixing put in place has grown to an estimated $40 million per year in the year 2000. The majority of work done to date has involved walls, either for excavation support, permanent retaining structures or hydraulic cut-off walls. The other major use for soil mixing has been for mass stabilization of soils whether for isolation of contaminants or for retaining structures. This paper will discuss the uses of and costs of typical soil mix structures. BACKGROUND Soil mixing is defined as the in-situ combining of reagents, typically cement, with native soils, the purpose being to modify or improve the ground. This history of soil mixing covers only the use of cement slurry, aggregated with soil in-situ by mechanical mixing methods only. Soil mixing by use of high-pressure hydraulic means, also known as Jet Grouting, will not be discussed. In addition this will not cover ex-situ (above ground) mixing, or mixing dry reagents with soil. Only mixing applications for building, civil and heavy construction applications will be discussed, with only passing reference made to uses of soil mixing for environmental remediation work.

Jan 1, 2001

The proper selection of a coal for a particular use depends on knowledge of fundamental data concerning the coals that offer, particularly their type and rank. Type of coal depends on the nature of the original plants and the environment in which they accumulated.Rank reflects the subsequent metamorphism of the coal, and is of two kinds-physical rank and chemical rank.INTRODUCTIONIn the different coalfields of the Commonwealth of Australia there occur many different kinds of coal which provide the coal-research worker with a wide range of material, but to the fuel technologist they represent complex problems in utilization.In the past it has been common practice to instal coal-using equipment of a design based on plant used in large industrial centres of other countries and then search for supplies of suitable coal. This has led to the transportation of coal mainly from New South Wales to other States, which have inadequate supplies of coal suitable for use in the equipment they have installed. However, modern economy and industrial practice demand local resources of coal for the development of industry, transport and domestic services, and a change in the old order is rapidly taking place. Technologists and engineers are now faced with the problem of designing and installing equipment to use whatever kind of coal is available in local coalfields.Much has been accomplished in this direction and important investigations are at present being carried out, but there still remains a wide field of problems for research.In planning and carrying out investigations into coal utilization it is essential that fundamental data and knowledge should be available and that its significance should be appreciated.

Jan 1, 1955

By A. F. van Weele

In general cast-in place piles will have an irregular pile shape. However, it is possible that these irregularities are so large that they may lead to pile deficiencies. It is very rarely possible to examine the pile foundations visually. Also the performance of a static or dynamic load test on each individual pile is not practical and much too expensive. This is the reason why pile integrity tests have been developed: to reduce the risk of inadequate pile foundations by performing a simple test. Pile integrity tests are non destructive tests which are carried out after a pile has been installed. Over the years a number of techniques have been developed of which the most important techniques are briefly described in this paper. Of all pile integrity testing methods, Sonic Integrity Testing is the most widely used method in western Europe. This method has been developed by TNO some 15 years ago. Since then the method has been improved, largely due to advances in micro-processing. A major improvement was made by the introduction of digital signal processing techniques. This made the Sonic Integrity Testing method more accurate (smaller defects can be detected), faster (more than 200 piles can be tested on a single working-day) and more reliable (the results are easier to interprete). A few cases will be presented which will show the usefulness of the Sonic Integrity Testing method. In one of these cases, a helpful tool for interpretation of the results is explained: the Sonic Integrity Testing Simulation program TNOWAVE. Notwithstanding the fact that Sonic Integrity Testing is a powerful method, it has also some strict limitations. These will be discussed at the end of the paper.

Jan 1, 1987