Search Documents

By I. G. Townsend, S. A. Buckley, C. I. House

"SummaryBritish Petroleum plc is developing a new gold recovery process known as coal gold agglomeration (CGA) for the recovery of gold from its ores. Liberated gold grains are recovered with high efficiency by adhering to agglomerates of coal and oil. The process is simple, can recover gold grains down to very fine sizes, avoids the large scale use of cyanide and provides the industry with an important new process option.A technical appraisal of the process has been performed on an extensive range of gold bearing materials. Ore types have been identified from which high gold recoveries can be achieved. A spreadsheet based computer program has been used to estimate the preliminary capital and operating costs of the process. Comparison with established gold production costs shows that the process costs of CGA, on the appropriate ore types, are potentially very attractive, which will enable economic treatment of lower grade ores than is currently possible using conventional cyanidation. Treatment of known placer deposits or gravity concentration plant tailings may be the most rapid means of demonstrating the process at a commercial scale prior to more widespread applications.1. INTRODUCTIONWestern world gold production for the 1980s has shown rapid growth sustained by a demand for gold at a favourable price (Fig. 1). Heap leaching and CIP/CIL operations have been responsible for the major part of new production during the 1970s and 1980s with present overall production costs averaging US$200/oz in Canada, USA and Australia and US$265/oz in South Africa [l]. A great deal of research and process development effort in the 1980s has been towards the treatment of refractory ores. Operating costs for the economic processing of these ore types by currently available routes are around US$300/oz [2]."

Jan 1, 1989

By S M Richardson, A W McNeill, J B. Gemmell

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. The Hellyer, Fossey, Que River and Mount Charter volcanic-hosted massive sulfide (VHMS) deposits occur within the Que-Hellyer volcanics of the Mount Read volcanics, Tasmania. These deposits occur along a 6 km strike length and are hosted by the same stratigraphic unit. We use a combination of whole-rock geochemistry, short wavelength infrared (SWIR) spectrometry, laser ablation inductively coupled plasma mass spectrometry (LA-ICP-MS) analyses and compiled historic whole-rock geochemical data to develop district-scale vectors towards VHMS mineralisation.Strongly sericite and K-feldspar altered rocks concentrate along a linear trend from Hellyer, through Que River, to Mount Charter. Ishikawa alteration index values and whole-rock major and trace element concentrations provide exploration vectors towards mineralisation on a regional scale as significant enrichment and depletion trends. On a deposit scale, composition variations in white mica (sericite alteration), chlorite and K-feldspar are observed and reflected in whole-rock geochemical variations. Regional SWIR data from drill core support the prevalence of phengitic mica and high-Mg chlorite within footwall alteration halos and highlight the main upflow zones of hydrothermal fluids in the VHMS district. The combined use of regional whole-rock geochemistry, SWIR and mineral chemistry is an effective way to highlight prospective areas for VHMS mineralisation in a mature exploration province.CITATION:Wu, S, Gemmell, J B, McNeill, A W and Richardson, S M, 2015. Exploration implications from variations in whole-rock and mineral chemistry around the volcanic-hosted massive sulfide deposits of the Que-Hellyer volcanics, Tasmania, Australia , in Proceedings PACRIM 2015 Congress, pp 85–90 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By Christopher Mark

A new method for the design of longwall pillars, called "Analysis of Longwall Pillar Stability" (ALPS), is described. The method is the final result of a research program that included two major field studies, extensive reanalysis of data from other in situ measurements, and detailed evaluations of available pillar design methods. The key problem addressed by ALPS is estimation of the abutment loads applied to longwall pillars. Based on underground stress measurements made in mines in three states, equations for predicting the magnitude, distribution, and time-of-arrival of abutment loads are proposed. Stress measurements are also used to evaluate empirical pillar strength formulas for use in longwall pillar design. ALPS has been verified by application to case histories from several mining areas, and examples taken from the Pittsburgh and the Blue Creek seams are presented. Step-by-step guidelines for using this design method are also given.

Jan 1, 1987

By M. Kuchta, S. Iverson

INTRODUCTION Underground blasting with perimeter damage control can reduce the requirement for loose rock scaling and reduce the risk of ground-fall injury. In previous work Johnson and associates (2010) from the National Institute for Occupational Safety and Health (NIOSH) developed a rock blasting perimeter damage and control design method included in the DRIFT software tool (NIOSH, 2018). This blast design software tool utilizes current methods of perimeter blasting with the added design element of a buffer row coordinated with the perimeter row to reduce perimeter damage. The coordination is in applying one of several blast damage models to calculate the damage extent for each production, buffer, and perimeter blast hole in the design. The calculated damage is then applied to the spacing between the blast holes. The basis of the buffer row design is the practical damage limit and is supported in this paper with computer models of signature blasts and full blast rounds using PFC2D (Itasca Consulting Group, Inc., 2018). Practical damage limit The practical damage radius (Rd) is not to be confused with burden. The practical damage extends toward the burden side and inward into the rock mass. For all practical purposes, the practical damage radius equals one half the burden (Hustrulid, 1999). The maximum extent of radial crack damage around a blast hole (Rcrack) and the maximum extent of radial crushing damage (Rcrush) are illustrated in Figure 1. A “practical” radius of damage (Rd) lies between the crushing limit and the cracking limit. The “practical” damage means that if the rock mass lying outside of the practical damage circle were removed, the rock remaining within the circle would easily break apart. Blast pressure and strain Work by Johnson (2010) and associates included signature blast tests in large concrete blocks. A signature blast test in concrete as exampled by strain data in Figure 2 was designed to monitor the dynamic strain at incremental distances from a single fully coupled blast hole, measure crushing and radial cracking extent by wire-sawing to expose the damage, and relate the damage to overbreak that is common in underground blasting. The initial slope or rising strain as the stress wave passes by the strain gages shows very high strain and strain rates close to the single blast hole. Since the derivative of strain versus time is strain rate, a line tangent to each of the strain curves at the maximum strain rate is drawn and the strain rates identified. The high strain rates were then plotted on a semi-log in Figure 3 to show a linear trend. The strain rate is greater than 100 s-1 within 20 cm from the blast hole for the signature blast experiment. The strain rate at the blast hole wall is projected to be 300 s-1. Figure 4 illustrates the damage in the NIOSH block exposed by a wire saw cut.

Jan 1, 2019

By M. W. Grabinsky, M. S. Diederichs, E. Hoek

Provided that they are used correctly, numerical models can be of considerable assistance to mining engineers in designing underground excavations. Guidance is given to the engineer or geologist who is not a specialist in numerical analysis but who is interested in using numerical modelling tools or in employing someone to analyse the stresses, displacements and failure in the rock surrounding underground excavations. A description of the theoretical background to a number of these numerical models and discussion of the advantages and disadvantages of each are followed by guidelines on the selection of appropriate models for each stage in the development of a mine, from exploration to the completion of mining

Jan 1, 1991

The Ngakuru zeolites are located about 20 km south of Rotorua and comprise a number of deposits dispersed along several fault lines. Relatively recent volcanic activity (250,000 years ago) resulted in ash depositions that vary in depth and mineralogy. Subsequent variable thermal alteration intensities generated predominantly clinoptilolite and mordenite zeolites with a range of physical characteristics. Typically the zeolites have low density (650 kg/m3), have high liquid (60% of dry weight) and odour absorption characteristics, and a high affinity for soluble cations (CEC =: 100 meq/100 g). Traditional kitty litter and oil absorbency markets are being expanded. New opportunities in waste and potable water treatment, organic and synthetic slow release fertilisers, use as a light weight aggregate for speciality concretes, and pozzalome activity are currently being researched.

Jan 1, 2000

By G. A. Bezuidenhout, S. Bradshaw, W. Malan

"The purpose of the study was to investigate the feasibility of SiC reduction of low-grade concentrates from Lonmin’s Rowland and Easterns operations with respect to metal fall and PGM recovery. These concentrates are rich in SiO2 and MgO with low concentrations of chalcopyrite and Cr2O3. Pd is the most abundant of the PGMs. SiC reduction of samples was conducted at 1600? with 2.5-3.5 kg SiC per 100 kg concentrate.PGM recoveries for Easterns concentrate were better than for Rowland. More than 85% of the Ir and Pd and almost 60% of the Pt were recovered with 3.5 kg SiC per 100 kg concentrate. SEM of slag samples showed little entrainment of metallic prills compared to Rowland samples. This was attributed to the relatively higher melt viscosities of the Rowland concentrate. In order to decrease slag viscosity and to enhance PGM recovery, the FeO content of the Easterns concentrate was increased with the addition of 10 kg converter slag per 100 kg concentrate. Ir and Pd recoveries were increased to about 95%, while Pt recovery was around 70%. On the basis of these results an optimum feed ratio between Easterns and Rowland concentrates and converter slag is proposed. Carbothermic reduction of the optimum charge was also compared to SiC reduction. Carbothermic reduction yielded a marginally higher metal fall; however, the calculated gas emissions and energy requirements were higher than for SiC reduction."

Jan 1, 2015

By D. Hoffmann, S. Plumb

IRUP is an iron-rich ultramafic pegmatite rock that formed due to hot ironrich fluids and gases replacing local stratigraphic zones of the Bushveld Complex. This study focuses on the estimation of the probability of IRUP occurrence on the Merensky Reef at the Marikana Karee Mine. A 2.15 million centare IRUP-rich domain at the Karee Mine was initially defined from the interpretation of a surface aeromagnetic anomaly, and exposures in the mine workings. Surface boreholes (spaced 250-500 m apart) within this IRUP domain contain approximately equal numbers of IRUP-replaced and IRUP-free intersections. Owing to the uncertainty in the continuity of the IRUP alteration (a result of the wide borehole spacing), the risk associated with the development of mining infrastructure within this domain is unquantifiable. Semi-quantitative data comprising visual estimates of IRUP replacement from reef development mapping data and surface borehole reef intersections was interpolated into blocks using indicator kriging estimation. Comparative analyses of the estimate of IRUP occurrences were made by changing the block size and declustering the data systematically. Reconciliations of the probability of IRUP predicted from the block models derived from the different data-sets and sequential indicator simulation models were analysed for four mining study blocks. A quantitative approach to modelling the occurrence of IRUP can be an additional tool for refining the estimate of the geological losses that inform the mine plan in such high-risk zones.

Jan 1, 2015

By G. W. Jones

One of the major activities of the Bureau of Mines is the promotion of safety in mining and other industries, In this connection, the Bureau determines the explosion hazards' of combustible gases, vapors, and solids and investigates means of controlling and preventing explosions. Many dangers are likely to confront men who must enter a building immediately after it has been damaged. Rescue workers must watch constantly for common gases and vapors that may be present in dangerous concentrations in and about demolished, buildings. Fuel gas is the most common source of danger, although refrigerants, gases from furnaces, flues, fires, sewers, and industrial processes, and gas from bomb, explosions in confined places also constitute definite hazards. To rescuers and trapped casualties, these gases offer the hazards of (1) explosion, (2) toxic effects, including asphyxiation and death, and (3) displacement of air so that the oxygen content of the atmosphere is reduced to or below the danger level.

Jan 1, 1944

By Derrick Dasenbrock

Two DFI committees independently organized two complimentary panel sessions for the recent American Society of Civil Engineers' (ASCE) Geo- Risk 2023 "Advances in Theory and Innovation in Practice" Conference in Arlington, Virginia, July 23-26. The Geo- Institute (GI) specialty conference welcomed more than 200 professionals and students interested in risk from around the world to the Washington, D.C., area for three full days of plenary keynote lectures, papers and no-paper sessions. The GI Risk Assessment and Management Technical Committee organized the event, which is the third in a series planned on a 6-year cycle. The previous two events were in 2011 in Atlanta and 2017 in Denver.

Sep 1, 2023

By M. Gomilanovic

The Republic of Monte Negro (a member of the Serbia and Monte Negro federation) has 26 different mineral raw materials, 4 of which are recoverable and 22 unrecoverable. At present, 12 types of mineral raw materials are exploited. Based on these resources more than 80 facilities were built for the production and processing of mineral raw materials. The most significant of them are coal, bauxite and decorative and technical stone. Bearing in mind that all the mentioned mining facilities have significant impact on the living environment, and that Monte Negro was declared the ecological state, the objective of this paper is to present the experience acquired from this field of science.

Jan 1, 2003

Several glacial episodes have been distinguished in the Shotover Valley. These produced fluvioglacial aggradational deposits downvalley from the terminii, commonly with well-preserved terrace treads, as exemplified near Skippers. Chronological links can be determined between the Shotover tills and sediments and assessed equivalent deposits in the Wakatipu Basin/Kawarau Valley region through axial valley profiling. Constructed relationships suggest older Shotover events were probably large and left little detrital deposits within the valley. In contrast, the younger advances were smaller and valley-constrained. They are modelled as being wholly of Waimea or Otira Glaciation age, advances from the latter period producing the visually distinctive flights of terraceremnants. Both the terraces and the inner gorge, now occupied by the Shotover River, have been extensively worked for alluvial gold. Enrichments in the former environment have predominantly been as bottom or intermediate lags produced during degradation phases of catchment erosion. Surface lags have been locally targeted. Placer mining in terrace areas involved both tunnelling and sluicing techniques. In the modem river the greatest concentrations were on the schist or false bottoms, although flood gold has always contributed to the returns. Placer mining techniques here were highly variable.

Jan 1, 1998

By S. Al-Thyabat

Conditioning of flotation feed is an important prerequisite for efficient flotation. This process depends on parameters such as conditioning time, slurry concentration and slurry pH. The objective of this paper is to optimize the conditioning process for the flotation feed. This was achieved by first conducting preliminary experiments at constant slurry pH to determine the significant ranges of slurry concentration and conditioning time. Then, factorial experimental design (32 experiments) was used to investigate the significance of the above parameters and their interactions on flotation recovery and grade. Statistical analyses using ANOVA approach showed that flotation recovery (R) and concentrate grade (g) were affected by feed conditioning. It was found that concentrate grade was linearly proportional to the slurry concentration and conditioning time, while flotation recovery was exponentially proportional to the conditioning parameters. A contour plot of the flotation response (recovery and grade) against conditioning parameters showed that the optimum operating parameters for flotation feed conditioning was a conditioning time of 1.5 to 2 minutes and a slurry conentration of 55% to 60%.

Jan 1, 2008

By Kosuke Hosoda, Shuji Owada, Hiroyuki Tokuichi, Hirobumi Ohta, Chiharu Tokoro, Moto-o Sato, Katsushi Okumura, Hiroto Ogawa, Hideaki Umezawa

The Japanese government enforced the "Small Domestic Appliances (SDAs) Recycling Law" in 2013. The law targets metals such as Ta, Nd, Dy, Co, and W to be recycled. Considerable amounts of these minor rare metals are mounted on printed circuit boards (PCBs) of the SDAs and/or e-scrap (electric and electronic scraps). Although precious metals and copper can be recovered from the e-scrap in non-ferrous metals smelters, most of the minor rare metals report to the slag in these smelters in non-recoverable forms. These metals are often used in particular types of devices, and therefore it is possible to concentrate them by separating such devices before feeding to the smelters. In previous work we developed technology for "Device Separation" to concentrate several minor rare metals, but some problems remained, such as how to remove the PCBs from e-scrap bodies, and how to detach devices from PCBs after their removal from escrap bodies. This paper proposes a two-stage crushing process in which, firstly, PCBs are removed from the e-scrap body with various crushers, and secondly, devices are detached from the PCBs involving the application of elevated temperatures. It was found that in the first stage, crushers with physically flexible rotating media such as chains, which provide high collision probability with the sample and consume relatively low energy, can quite effectively separate the PCBs from the e-scrap bodies non-destructively. In the second stage, heating was found to be remarkably effective in helping to detach mounted devices from PCBs, also non-destructively. This paper will also discuss the detailed mechanism of the heating effect on the detachment of devices from PCBs.

Jan 1, 2014

By Christopher J. Bise

INTRODUCTION Traditionally, the purpose of mine ventilation has been defined as diluting, rendering harmless, and sweeping away hazardous gases which may accumulate in the mine workings. In addition, the ventilation sys- tem must keep the mine air as close as possible to the composition of the outside air in order to provide a healthy environment for mineworkers and to prolong the useful life of mine equipment. When plans are being developed for opening a new mine, sound engineering dictates that a calculated estimate be made of both the maximum air quantity requirement and the accept- able mine pressure (head) loss related to the under- ground layout and projections. For efficient mine ventilation, the two problem areas that demand constant attention are (1) the control of fugitive air (through short circuiting or leakage) and (2) the provision of ample airway capacity. Fugitive air is that quantity of air which enters the mine but short-circuits the mine ventilation system through improperly constructed stoppings, doors, overcasts, etc.; it represents an investment in power consumption (to move the air by the fan) while providing little, if any, return on the investment. Typically, efficient mines are those that can deliver 50% of the air handled by their fans to the working faces. Paying close attention to a new mine's ventilation system at the outset can provide numerous benefits during the later stages of mine development. It has often been noted that the main cause of trouble in many mine ventilation systems can be found within a radius of 2500 ft from the fan installation. Ample airway capacity is essential to maintain the mine's head loss within manageable limits. Similar to an industrial ventilation setting, the mine's headings are the ductwork for the handling of air. The difference between normal industrial ventilation and mine ventilation is apparent when more capacity is needed; a plant can install new or larger ductwork, but it is difficult for a mine to install new entries or enlarge existing entries. Air velocities are usually maintained under 800 fpm to limit the stirring up of dust in the intakes. With this in mind, it is usually simple to estimate the number of entries required in the ventilation system. The problem that often arises, though, is when the system's capacity is overestimated and the required head loss to achieve the desired quantity is excessive. Experience has shown that problems are often encountered when the mine's condition requires much more than three inches of water gage to ventilate the mine on a day-to-day basis. Thus, if proper attention is not paid to these parameters in the preplanning and initial development of a new mine, it may take an extremely high water gage to achieve the ultimate quantity. This chapter will not delve deeply into the theory of mine ventilation; many other books have been writ- ten on that subject. Instead, the following sections will point out the important considerations that should take place before the development of a mine ventilation system, as they relate to normal industry practice. FUNDAMENTALS OF AIRFLOW Volume Versus Water Gage The common method of measuring ventilation pressure is in terms of inches of water gage (wg); one inch of water corresponds to a pressure of 5.2 psf (pounds per square foot). Further, airflow follows a square-law relationship between volumes and pressures where, as shown in the previous section, a doubling of quantity requires a quadrupling of pressure. Table 1 shows how heads and quantities are related. Using Table 1, it is readily apparent that, for the same mining conditions, an increase in pressure from one inch wg to three inches wg results in only 34,640 cfm when the initial quantity is 20,000 cfm. Pressure Losses Pressure losses in mine ventilation systems are caused by friction between the heading walls and the air stream and by shock losses due to abrupt changes of the air stream caused by obstructions. The common pressure-loss formula for mine ventilation is: [ ] where H is the pressure loss in inches of water gage, K is the coefficient of friction, L is the length of the airway in feet, 0 is the airway perimeter in feet, V is the air velocity in feet per minute, and A is the cross-sectional area of the airway in square feet. Table 2 lists the US Bureau of Mines schedule of friction factors (K) for mine airways. Recent research indicates that the values of K are slightly less for coal mines than those presented in the USBM Schedule, and Table 3 should be consulted in this instance. It should be

Jan 1, 1986

By Dave Bulled, Khiratt Husain

"The use of High Pressure Grinding Rolls is gaining significant interest as a replacement for SAG milling in the comminution of hard rock ores because of the potential for reduced costs in terms of energy and grinding ball consumption. Estimation of expected energy consumption for fullscale operations has conventionally involved the testing of a few fairly large samples in pilot scale HPGRs. This paper describes the development of a small-scale batch test method for determining a work index for HPGR of a large number of samples that can be used to show the variability of the ore as an integral part of a geometallurgical approach to circuit design and production forecasting.The test equipment that was developed comprises a hydraulic press with adjustable pressure operating on small batches of ore in a locked cycle test procedure. A work index measurement was determined that links the specific energy consumed in the test to the breakage achieved such that the index can be used to predict comminution energy requirements.The measurements were calibrated against laboratory scale HPGR equipment to develop a model for design and forecasting that includes crushers, HPGRs and ball mills in a complete comminution circuit.INTRODUCTIONEstimation of the expected energy consumption for full-scale HPGR operations has conventionally involved the testing of a few fairly large samples in laboratory and pilot scale machines (Danilkewich, 2006). With the financial input of five mining corporations, SGS Lakefield Research has developed a small scale test capable of providing relevant information on the variability of energy consumption and size reduction from drill core samples. It is our intention that the values measured in the test will be scaled up, through benchmark comparison with pilot sized and full sized HPGR units, to allow plant design and throughput forecast on a block by block basis of the mine plan using a new CEET program. CEET® (an acronym for Comminution Economic Evaluation Tool) is an advanced grinding design and performance predicting tool that uses energy relationships to model the performance of SAG / Ball mill circuits."

Jan 1, 2008

By C. Rutledge

The new mining method of multiple strata blasting, called StratablastTM, is described in a separate paper presented at this conference. The current paper presents three case studies where the technique was implemented in three different configurations in large-scale blasts in Australia. The first two blasts were in the Hunter Valley of NSW while the third was at Ensham Mine in Central Queensland. All blasts comprised multiple layers that were blasted in a single drill and blast cycle, with unique designs applied to each layer to achieve different blast outcomes from the various layers.

Jan 1, 2006

By S Piercey, M Doublier, B Eglington, D L. Huston, S Pehrsson

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. The distribution of volcanic-hosted massive sulfide deposits through time is episodic, involving relatively short time intervals of apparent high productivity and long intervals of low productivity. This distribution is principally related to the geodynamic evolution of the Earth. The vast majority of these deposits is associated with the assembly of supercontinents and form along convergent margins, generally in back-arc basins or rifted arcs. This contrasts with the distribution of black smoker deposits, which presently form along both convergent margins, particularly in the western Pacific, and divergent margins at mid-ocean ridges. However, the latter environment is rarely preserved deep into the geological past.There also appear to be systematic patterns in a number of other features of volcanic-hosted massive deposits through time. Deposits hosted by mafic-dominated successions are dominant in Archean to Proterozoic rocks, whereas deposits hosted by felsic-dominated successions are most common in Phanerozoic rocks. Geochemical characteristics of felsic rocks within the host successions and ore lead and sulfur isotope data of the ores have also changed with time. This suggests changes in the process of subduction and/or changes in the characteristics of the over-riding plate with time, consistent with isotopic and geologic evidence for greater reworking of pre-existing crust during the Phanerozoic.Other secular changes that are reflected in characteristics of volcanic-hosted massive sulfide deposits include the redox state and salinity of coeval oceans. As an example, both the abundance of sulfate minerals and sulfur isotope patterns reflect the development of periods of anoxic conditions over time, with major intervals of anoxia reflected by a virtual lack or low abundance of sulfate minerals in volcanic-hosted massive sulfide deposits in the Mesoarchean to the Palaeoproterozoic, and shorter intervals since.CITATION:Huston, D L, Eglington, B, Pehrsson, S, Piercey, S and Doublier, M, 2015. Secular variations in the abundance and characteristics of volcanic-hosted massive sulfide deposits – implications for evolution of convergent tectonics and ocean geochemistry , in Proceedings PACRIM 2015 Congress, pp 475–480 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By Georgene Renner, William R. Yernberg, Steve Kral

ABB provides integrated pro-cess control, optimization and information systems, drives and power-distribution equipment as well as other services and support to the mining industry. ABB also supplies electrical equipment and control systems for the openpit mining industry. This includes products for the automation of electrical equipment in mine shovels. The company offers low- and medium-voltage drives for various mineral- industry applications. And ABB provides gearless mill drives for equipment, such as semiautogenous grinding mills that are too large for conventional drives. ABB Industrial IT provides innovative hardware and software systems that cover all the requirements of the modern mining and processing plant.

Jan 1, 2004

By G. Belardi, N. Shehu, L. Passeri

"This paper illustrates the possibilities of applying what is essentially a physical soil-washing technique for treatment of soil contaminated by semi-volatile hydrocarbons in order to reduce the level of pollution to that permitted by Italian regulations.The ground around a factory producing lubricating-oil additives consists of silty-sandy soils which have long been contaminated by heavy hydrocarbons, especially branched alkyl benzenes and naphthalene derivatives. Pollution is particularly marked in some zones (hot-spots) where Total Petroleum Hydrocarbon (TPH) values are as high as 7000 ppm. The hydrogeological characterization showed the presence of a gravely soil in a sandy/silty matrix, with big stones making preferential paths to the water circulation. The permeability is 10-5 m/s and the water-table occurs at about 3 m below ground level.After laboratory and pilot-scale studies on representative samples of in-situ soils with TPH between 2700 and 7000 ppm a flow-sheet is proposed for reducing TPH values below the acceptable limits in Italy, namely 700 ppm for hydrocarbons with more than 12 carbon atoms and less than 250 ppm for those with fewer.On the basis of indications provided by laboratory tests a mainly physical treatment process of soil washing has been developed. This is used to clean-up soil from a site polluted by non-volatile hydrocarbons with a high TPH value. The process involves a high-density attrition stage followed by wet screening at 0.5 mm and then wet-gravity separation of the minus 0.5 mm fraction in the centrifugal field.The gravity separation products consist of a light fraction with particle size finer than 15 micron which contains most of the pollutants, and a heavy coarse fraction with TPH values conforming to the limits set by Italian regulations (less than 700 ppm hydrocarbons whose structure contains more than 12 C atoms, and less than 250 ppm of hydrocarbons with less than 12 C atoms)."

Jan 1, 2003