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By Dirk Timmerman, Klaus Schönert

This paper reports on the results of research on counter flow classification of - 10 µm particles. Two classifiers were designed and constructed, the first semicontinuous and the second fully continuous. The latest version operates continuously with recoveries of fines up to 90% at feed loads of 30 vol.%, separation sharpness X25/x75 up to 0.83, without bypass and cut sizes down to 1.4 µm (x50 from separation efficiency curve).

Jan 1, 1995

By J. Raatikainen, I. Auranen

"INTRODUCTION Sorting technology has advanced rapidly recently. Of the two possibilities particle sorting has advanced especially in introducing new sensors and the use of multiple sensors. The analyzing and computing lead times have also shortened. This has made it a viable technology option for many applications but not for the most important sorting task at a mine; the run-of mine ore. Bulk sorting, however, has advanced in smaller steps, even if the technology has the capability to tackle the run-of-mine ore and offers a high potential to bring mines substantial economic benefits. The sensors, which are useful are restricted by the difficult operating conditions or by the ore characteristics. One of the important features of run-of-mine sorting technology is the major material handling infrastructure required. The capital expenditure can be substantial requiring good understanding of the sortability curve (grade-recovery curve) behavior of the ore. This paper discusses one industrial case, where substantial amount of data measured from the ore feed belt has been obtained. This data was then analyzed for the grade recovery curves obtainable from that ore with sorting. EXPERIMENTAL The XRF measurement used was the FCA device from IMA Engineering (Fig.1). It is a specially built version of XRF designed to operate with coarse material. The measurement was situated at the main belt conveyor taking ore from the primary crusher to the stockpile. The crushed ore was about 80% -300 mm at a capacity of 1100-1200 tph (Fig.2). Measurement times were varied from 30 seconds to a minute. Measurements were obtained over several months.As the XRF is a surface distribution measuring technique, the measurement was first tested upon accuracy and reproducibility against plant daily wet chemical assays and against on-line Courier feed analysis. As there is a varying delay in the ore stockpile, the delay time was used as a variant in the validation. Different measurement times were tested for accuracy. A one minute measurement time gave a total measurement relative standard error, which was both for Cu and Ni below 3%. The results showed also very robust correlations; for Cu, the correlation coefficient was R > 0,94 for one minute measurements over three months and R > 0,86 for Ni respectively. Fig 3. Shows a measurement period of two hours showing substantial feed grade variations (given as relative value; 1.0 represents the mean value during the measurement month)."

Jan 1, 2018

By Thomas L. Hurst, Joseph M. Hans

INTRODUCTION The former Shiprock uranium mill site is located in the community of Shiprock, New Mexico on a bench overlooking the San Juan River (Figure 1). Both Shiprock and the mill site are located on the Navajo Indian Reservation. The millsite was built on a 230 acre tract leased by Kerr-McGee Oil Industries Inc., and was operated from 1954 to 1963. The Vanadium Corporation of America and its successor, the Foote Mineral Company, operated the mill from 1963 to 1968. During its entire operation, 1.5 million tons of uranium ore was milled producing 3711 tons of U308 concentrate. The concentrate was sold to the former U.S. Atomic Energy Commission. During milling, the ore was crushed to less than 35 mesh and acid leached to remove the uranium. The [ ] uranium was removed from the leach solution by solvent extraction and the uranium-barren solution discharged into raffinate ponds for disposal by evaporation and seepage. The spent solids (tailings) were pulped with waste process fluids and water and discharged to form tailings piles. The piles (ponds or dams) were formed with earth starter dikes. The dike walls were then raised using the separated sand fraction of the tailings. The slime fraction of the tailings were discharged into the diked impoundment. Since only a small fraction (0.25 percent) of the ore was U308 by weight, most of the ore bulk ended up as wastes. The wastes contain nearly all of the uranium decay chain radionuclides present in the ore except uranium. The mill was shut down and partially dismantled before the lease expired in 1973, and the remainder of the mill and wastes reverted back to the Navajo Tribe which was the lessor. Portions of the millsite were then occupied by the Navajo Engineering and Construction Authority (NECA), a Tribal enterprise. In order to provide skilled personnel to operate earth moving equipment, a heavy equipment operator's training school was also established on the mill site and was administered and operated by NECA. The U.S. Public Health Service (PHS) and the Environmental Protection Agency (EPA) staff conducted radiation surveys at four inactive uranium millsites on the Navajo Reservation in April 1974. During the survey of the Shiprock millsite, it was noted that the training activities were degrading its radiological condition. Both the PHS and EPA staff recommended that the training activities be directed towards the decontamination of the millsite area and interim stabilization of the tailings piles. Preparation for the decontami-nation work commenced in April 1974. In order to facilitate the decontamination activities, the mill site and its contaminated environs were divided into major areas labeled A through F (Figure 2). Areas A, B, and C are defined as "perimeter" areas and contained surface contamination ranging from 5 to 30 cm deep. Area D defined the mill yard and contained contamination ranging from 0.6 to 5.5 m deep. Area E contained several raffinate ponds which were removed to line up the northwest boundary of the tailings pile and to move the contaminants away from the vicinity of the occupied on-site buildings. The contaminants in Area E ranged from 0.05 to 5.5 m in depth. Area F encompassed the ore storage yard. Contaminants in this area ranged from 0.18 to 1.2 m deep. Washes, however, had contamination exceeding 7.6 m deep.

Jan 1, 1981

By Weixiang Meng, Zhao Song, Ming Guo, Zhigao Liu, Shoubao Zhang

"There are many multi-layer coal reserves in China which account for nearly 20 to 25% of the existing proven reserves. When the coals are mined in a descending order using the fully mechanized long wall mining method, they usually have roof collapse with cavity and air leak in roadways and long wall stopes; this is especially relevant when the thickness of the rock between two coals is less than 5.0 m, as the roof stability of lower seam would be greatly reduced. At the Wuhushan Mine, such abnormal strata behavior occurs, including abnormal support stress, irregular period roof pressure, and roof falls in the face. The main reasons for these behaviors include smaller roof thickness, lower rock strength, and mining influence of upper coal; these reasons were determined by analyzing the in situ conditions. A mechanics-based model of the mine geometry was set up to study the sources of stress, the stress evolution of coal, and the discriminant method of roof stability; to obtain the smallest safe thickness of roof between two coal layers, numerical modeling was applied to analyze the relations between the roof thickness and its displacement distribution. Based on the above research, the No.1001 face was divided into easy roof caving zone (ERCZ), roof caving transition zone (RCTZ), and unlikely to roof caving zone (URCZ). In the ERCZ, comprehensive grouting to reinforce the roof was implemented; in the RCTZ, local grouting to reinforce the roof was implemented. Furthermore, a face stability control index and its range were developed in a fully mechanized face combined with mine pressure observation. After implementing this strategy, there have been no unexpected roof caving or other serious strata behaviors.INTRODUCTIONChina is the country with the largest coal production in the world, and the conditions for its coal seams are particularly complicated. For example, in China’s Datong, Wuda, Pingdingshan, and Qinshui mining areas, there are a large number of multi-layer coal seams. The thickness of these coal seams ranges from 1.3 m to more than 10 m, and the spacing between layers is from under 1.0 m to over 30 m. A common problem with these coal seams is that one layer of mining has a large impact on other coal seams, which are customarily referred to as close-distance coal seams. According to statistics, close-distance coal seams account for about 20–25% and 15–18% of China’s coal reserves and production, respectively.From the reference research of Huang (1999, 2002, 2018) and Tu (2010, 2011) we can learn that in the mining process of close-distance coal seams, downward mining is mainly used. When the upper-layer coal is mined, it has a huge mining disturbance effect on the floor (that is, the roof of the lower coal). The mining disturbance increases the mining fissure in the rockmass and weakens the strata strength. The mining problem formed by this issue is that when the lower coal is mined, problems such as difficulty in roadway support, collapse of stope roof, and difficulty in ventilation are often encountered. Especially for the mining of medium-thick coal seam groups with a layer spacing of less than 5.0 m, there has been a large-scale roof collapse accident with a fall height of more than 10.0 m. In response to such problems, domestic and foreign scholars and engineers such as Tan, Zhao, and Xiao (2010), Xie et al. (2017), and others have also carried out targeted theoretical and practical research."

Jan 1, 2019

By L. Alan Weakly

The Powder River Basin of northeastern Wyoming and southcentral Montana contains the largest strip minable reserve of compliance coal (0.27 kg or 0.6 or less Ibs of sulfur per million Btu) in the United States. In mining parlance, it is the mother lode. Wyoming's portion of the Basin is called the Southern Powder River Basin (SPRB). It has been the nation's top coal producer for the past nine years. And based on 1997 production statistics and estimated strippable compliance coal reserves, Wyoming should remain the nation's top producer for the next 51 plus years. In August 1994, ME featured 10 articles that covered all aspects of coal production in the SPRB. In this issue, five authors update the 1994 articles and provide a pre- diction of what the Basin will look like at the turn of the century. In addition to this article, the titles and authors of the other five articles include: ? "Geology of Wyoming's Powder River Basin coal field," by Gary B. Glass and Robert M. Lyman. ? "Coal quality challenges in the Powder River Basin," by Sharon Roness, Richard A. Turpin and Robert A. Young. ? "Markets for Southern Powder River Basin coal," by , Patti J. Best and Patrick W. Wright. ? "Evolution of mining equipment in the Powder River Basin," by Kenneth R. Miller and James L. Long. ? "Hauling coal into the 21st century," by Dave Willis and Will Cunningham.

Jan 1, 1998

By M. E. Lewellyn

A new class of thionocarbamates, O-alkyl-N-allyl thionocarbamate (III), has been found to give better metallurgy than 0-isopropyl-N-ethyl thionocarbamate (I) for a number of sulfide ores containing copper, lead, and zinc. Data on the results of flotation tests with several such ores are presented. The improved metallurgy is believed to be due to the participation of the allyl double bond in the complex formation of the thionocarbamate on the surface of the mineral.

Jan 1, 1985

By Trevor J. Smith, Kevork A. Chouzadjian, Rob S. Burns, Steve J. Roden, Michelle A. Spieth

Investigations were undertaken to improve the recovery of silver from CRA's Hidden Valley gold/silver deposit by the addition of calcium peroxide. The addition of 0.75 to 1.5 kg/mt of calcium peroxide improved silver extraction by 5 - 300 and reduced cyanide consumption by up to 30%. Gold extraction was also marginally improved in some of the tests. The results are of economic interest and plant trials will be initiated once operations are underway.

Jan 1, 1990

By M. T. Ityokumbul

Experiments have been carried out to determine the efficacy of bubble column flocculation. An acid-base neutralization method was used to determine the mixing time in a 0.064 m x l.5 m bubble column equipped with a Mott porous sparger. The mixing times decreased as the tracer injection point approached the gas sparger. Preliminary flocculation experiments on two clay samples were carried out. The effects of frother addition and air flow rate are presented and discussed.

Jan 1, 1998

By R. D. Rein

The Henderson mine is a large continuous panel caving operation in central Colorado. The orebody is roughly 760 m by 910 m and will be extracted from two production levels. Each level is divided into three north - south panels 244 m to 317 m wide. The general arrangement of a production level and its associated undercut drifts is shown in Figure 1. The production drifts (3.6 m by 3.7 m) are spaced 24 m apart center to center and have undercut drifts (3.4 m by 3.8 m) 17 m directly above them. The production drifts are interconnected with drawpoint crosscuts (4 m by 4.3 m) driven on 12 m centers. Drawpoint crosscuts receive substantial brow steel and concrete support prior to undercutting. The brow steel is a 15 cm by 46 cm wide flange beam with a 2.5 cm wear plate to box in the 'cave' side of the beam. The concrete is a minimum of 0.3 m thick and the batch design results in 27 500 kPa strength. The finished drawpoint opening is 3.7 m by 3.2 m with a choke area of 3.7 m by 2.4 m. (Figure la) Undercutting and bell formation is accomplished by blasting rings of 7.6 cm holes drilled from the undercut drifts. The rings have a burden of 2.0 m and six rings develop one complete bell.

Jan 1, 1985

By Regis F. Kopac

With the growing dependence on coal as an energy source, finding a means to prevent or reduce the freezing of coal during winter shipment has become increasingly important. Several chemical companies focused on this problem and applied chemical technology to develop suitable solutions. As a result, a number of freeze conditioning agents have been introduced in the past few years, several of which have been highly effective in controlling frozen coal. The efficacy of these chemicals has been well established, however many users have found that only a well defined program will produce acceptable results at a reasonable cost. Positive program results yielding recognizable benefits have caused many power generating and coking plants to consider freeze conditioning agents. A freeze conditioning agent (or FCA) has been defined as a multi-component organic liquid designed to be sprayed on coal particles at bulk shipping points to assure that receiving locations will be able to unload, store and use the treated coal with minimum delay or inconvenience. The chemicals are generally applied at rates of 1 to 2 pints (0.473 to 0.946 liters) per ton and normally function by changing the crystalline structure of water left on the coal during processing. The bond strength of the ice that may subsequently form is markedly reduced. Instead of freezing into solid, unmanageable masses, the coal can be easily unloaded by rotary dumping or the opening of hopper car doors.

Jan 1, 1980

By Wenbing Guo, Mingjie Guo, Hebing Zhang

As resources are depleted in coal mines, isolated support pillars are often mined. Cemented paste backfill has been used to provide support during the mining of these pillars due to its safe and green mining characteristics. However, traditional cemented paste backfill mining has some shortcomings, such as poor adaptability to geological structure, leakage of paste during filling, imperfect roof-contact in partially backfilled areas, and insufficient roof strata support before the backfill reaches its design strength. This paper presents the results of a case study where L-shaped interval cemented paste backfill technology has been safely used to mine an isolated coal panel beneath existing surface structures. This backfilling method has the advantages of preventing paste leakage, improving the backfilled ratio, reducing surface subsidence, and strongly adapting to geological structure. During this field application, the setting pressure of the backfill shield reached 20 MPa, the compressive strength of the backfilling body reached 2.0 MPa, and the backfilled ratio reached 93.7%. Observational ground monitoring showed that roof-to-floor convergence in adjacent roadways ranged from 132 to 167 mm. No obvious fractures or separations in the overlying strata were observed by a borescope. The maximum surface subsidence was 342 mm, the subsidence factor was 0.136, and the damage to surface buildings was within Grade I (slight damage). This case study clearly demonstrates that the L-shaped interval cemented paste backfill method can be safely used to mine the isolated coal panel beneath surface buildings.

Jul 1, 2023

By R. I. Yano

Analyses of trace elements in bulk ores and in chalcopyrite and other ore minerals from porphyry copper systems worldwide, including skarns and breccia pipes, reveal some key factors regarding byproduct critical elements. Petrographic and scanning electron microscope analyses combined with a laser ablation inductively coupled plasma mass spectrometry technique developed by the USGS on selected ores from the Mackay-Stanford Ore Deposits Collection suggest that selenium occurs as atomic substitutions in sulfides, particularly pyrrhotite (up to 430 ppm) and chalcopyrite (up to 300 ppm Se); tellurium occurs predominantly as less-than-five-micron-sized inclusions of bismuth-tellurium minerals in other ore minerals, particularly chalcopyrite; cobalt and nickel are locally enriched in pyrite (up to 7800 ppm Co and 9100 ppm Ni); gallium is locally enriched in magnetite (up to 83 ppm Ga); indium and silver are locally enriched in chalcopyrite (up to 890 ppm In and 770 ppm Ag); cadmium substitutes for zinc, particularly in sphalerite (up to 4100 ppm Cd). There is considerable variation, both between and within individual deposits, in both selenium and tellurium content in these copper ores, ranging from 0.6 to 148 ppm Se and 0.09 to 64 ppm Te. Indium tends to be high in copper ores that are also enriched in silver. At current prices, many of these elements are sufficiently enriched to warrant investigation into metallurgical recovery as byproducts of the major elements.

Feb 27, 2013

By Stanislaw Siewierski

One of the major problems in the exploitation of deep copper ore deposits in the Legnicko-Glogowski Copper Field is ground control and, in particular, roof control of the headings. The depth of copper deposits ranges from 600 to 1200 m. The parameters of ore-bearing rocks and the surrounding rocks show considerable variance of strength properties along the mining thickness (Fig. 1). The roof of the openings exploitation headings consists of dolomite-calciferous rocks. The roof strata consist of layers, the thickness of which vary from several centimeters to 0.6 m (sometimes to 1.0 m). These strata differ in strength and are characterized by igneous and sedimentary discontinuities. On the basis of studies in structure, properties and geological mining conditions of the roof strata in Polish copper mines, the method of roof classification has been developed. (1) This method uses the stability index "L" of the roof, calculated from [ ] The value of "t" ranges from 0, l to 0,8. Based on the value of "L", four classes of the roof (Table 1) corresponding to different degrees of fall hazard have been developed. The mining of copper ore deposit is carried out by a single lift room-and-pillar mining system or a two-stage room-and-pillar system with roof caving. (3, 4). The single lift room-and-pillar mining system is most often used. It consists in working of the deposit of 2.5 to 4.5 m mining thickness, 150-350 m long mining line, consisting of parallel room faces 6 m wide with pillars 5.7 or 16 m wide. As the mining line advances, cross-cuts joining individual rooms are made, thus forming support pillars 3x5, 7x7, 5x16, and 7x16 m (Fig. 2). The two-stage system (Fig. 3) mines the deposit with rooms 6 m wide into blocks 25 x 35 x 25 x 45 m. During the second step these pillars are worked with rooms and cross-cuts into pillars about 3 x 4 to 4 x 5 m (retreat mining). The pillars are crushed just before the fall of the roof. The maximum opening of workings in the two-step system is about 10 m, i.e. one belt entry and one row of bearing pillars. In single lift room-and-pillar systems it is about 35 m., i.e. three belt entries and three rows of bearing pillars. The support of the headings consists of expansion bolts 1.6-1.8 m installed in a bolting pattern 1.1 x 1.5 m. In both systems the fall step is about 10 m, i.e. the width of a belt and a bearing pillar. The Indications of Rock Mass Pressure Individual Dog Heading Individual dog headings are set wide apart from each other in the unmined ore deposit area where there is no interaction of headings. The indication of rock mass pressure are, in such cases, limited to the deformations and damage of side walls. Radiometric studies of rock mass loosening showed that the maximum range of lifts in side-walls immediately after web was 2 m (4,5). In the course of time the cracked side rocks fall off, thus widening the heading. The morphology of side-walls is influenced by the location of the heading in relation to clump layer/tarry. This layer is characterized by a weak strength and "lubricating" properties. The location of the heading in

Jan 1, 1982

By S. K. Kawatra, S. J. Ripke

Iron ore pellets use bentonite clay as binder at dosages from 0.5to 1.0% (10 to 20 pounds per ton) of moist iron ore concentrate. Bentonite is typically shipped hundreds of miles from Wyoming to the Lake Superior iron ore district. It is therefore best to know how effective a bentonite will be before it is shipped and used to bind concentrate into pellets. Current tests of bentonite effectiveness include: plate water absorption, methylene blue uptake, free swell, exchangeable cations, and glycolated layer expansion. However, these tests have many disadvantages: First, they only measure the ability of a bentonite to expand, disperse, or absorb, not how well it will bond. Second, most of the tests require up to 20 hours to perform. Bentonite only has a few minutes to interact within the concentrate before it is indurated (sintered). These long-term tests may not reflect the true behavior of bentonite. Third, the tests are tedious and often results are inconsistent between different iron ore concentrates, labs, and technicians. Fourth, some of the testing equipment is difficult to procure. Fifth, procedures are often difficult to standardize. In this paper, the Binder Effectiveness Strength (BESt) test to evaluate binders for iron ore pellets is presented that: measures the ability of bentonite to bond, requires only about two hours to complete, is simple and reproducible, and utilizes readily available equipment that has been standardized for decades.

Jan 1, 2003

By Louis W. Cope

Placer material, regardless of the value of gold, tin, titanium, diamonds or other value mineral it contains, remains worthless until the valuable constituent is separated from the bulk of the material. This separation is no mean task when considering that a gold placer containing 0.01 to 0.03 oz/yd3 of gold, which is the range of grade of typical placer deposits; yes, 1:3,000,000! It is a miracle that gold placer mining has ever been profitable. Added to the situation of minute quantities of gold are several other problems, such as the gold often being very finely divided, the association of boulders and clay, and running and/or ground water. As each placer deposit is individual and distinctive, many factors have to be considered in planning the method for gold recovery from the alluvial deposit. Previous chapters have shown that in the exploration and sampling stage, some knowledge of the amount of each size of coarse and fine material which occurs in the deposit as well as the size distribution of the discrete gold particles is determined. Unless informed to the contrary, as a general rule of thumb, plant designers will consider that 40 percent of the material will pass through a 5/8-inch screen or punched plate hole. Also that there may be an occasional nugget and normally considerable amounts of fine gold. The designer must

Jan 1, 1992

By Max John

INTRODUCTION One of the main advantages of the New Austrian Tunneling Method (NATM) is its adaptability to varying rock conditions. This method is therefore successfully applied when complex geologic features are expected which cause uncertainties in the prognosis of the rock mass behaviour, as was the care at the Arlberg Expressway Tunnel. Furthermore, by employing the NATM problems are avoided when encountering varying geologic conditions or even extension fault zones along the tunnel, as was the case at the Dalaaser Tunnel and Perjen Tunnel. The NATM can also be used for rock and soil. The following examples will show how various rock conditions can be mastered. In accordance with the NATM all tunnels which will be discussed are road tunnels having a cross section area between 70 and 100 m2. The construction consists of a primary support, which ensures stress redistribution during excavation (Figure 1). The primary support is reinforced shotcrete strengthened by steel arches, when necessary, and rock bolts. When encountering rock pressure, swelling rock or soil, the primary support is suplemented by an invert arch. The purpose of the secondary lining is to increase the safety of the tunnel with regard to long-term stress redistribution processes or changes in rock mass properties as well as to protect the insulation. The secondary lining consists of cast concrete without reinforcement and is generally 0.25 or 0.30 m thick.

Jan 1, 1981

By Gerald L. Colborne

The major role of financial institutions in the free world in the development of mineral resources is to provide the necessary capital in the form of loans to underwrite the development of these resources by the industry on a stable and continuous basis. This capital provided by the depositors to the bank must be invested in sound projects where the degree of risk is relatively small, the return on the investment is reasonable and the time to retire the loan is of short duration. Money is, therefore, one of the important ingredients along with technology, manpower and an exploitable resource that will provide a sound economic base in this region of development. Canadian banks are similar to banks of other western nations, although each country has some differences in their banking systems. The basic difference between the American and Canadian systems is that the former employs unit banking while the Canadian utilizes the more mobile branch system, which is prohibited or restricted in the United States and some other countries. Canadian banks are the subject of this paper, so, subsequent material will relate to Canadian banks unless indicated otherwise. It will be realized by everyone that the mineral industry is only one area of the economy in which the banks are interested and to cover the subject matter in the time allowed, the topic is restricted to mineral resource development. In Canada this is a vital sector of our economy and before we outline some of the details of financing and services involved in various aspects of mineral development, a review of the Canadian mineral industry is a necessary part of the sequence of presentation. The writer presents this material as a resource engineer advising a major bank in mineral matters, so the presentation may relate more to the financial problems of the developer

Jan 1, 1972

By M. Gaft, G. Panczer

Sedimentary francolite and magmatic apatite are characterized by specific laser-induced time- resolved luminescence, which allows phosphate to be distinguished from accompanying dolomite. The "selective window" parameters are: he, = 337 nm, delay time after laser pulse is 0.5-5 ps, and detection spectral range is 400-525 nm. In the case of nonuniform distribution of phosphate minerals, it makes it possible to accomplish lasero- luminescent sorting and to receive intermediate product with a lowered dolomite contents.

Jan 1, 1999

By Edward F. Divers, J. Drew Potts

Powered dust-filtering helmets (dust helmets) have become increasingly prevalent in the underground mine environment in recent years. Many longwall operations use dust helmets to reduce the respirable dust exposures of miners required to work downwind of crusher, roof support and shearer operations. While continuous mining sections are usually less dusty, use of diesel equipment represents a new potential health problem. The U.S. Bureau of Mines recently conducted a study to determine the efficiency of a dust helmet on diesel-size particulate. Two underground evaluations were conducted on continuous mining sections that were using diesel haulage vehicles. Model No. AH21 Racal airstream helmets equipped with AS22-3 prefilters, AS23-3 main filters and AS60476 tyvek full-face seals were tested. Helmets were placed on mannequin heads and mounted in the ram car cabs. Andersen Model 298 (eight-stage) personal cascade impactors were used to sample air inside and outside of the helmet. Impactor stages 8 and the final filters were analyzed to determine submicrometer (0 - 0.93 µm) particle concentrations. Results from this study showed that the dust helmet was 94% (± 3% at 95% confidence level) efficient on submicrometer particles. Typically, very little coal dust is observed in the submicrometer size range, therefore, these particles were assumed to be from diesel exhaust. Results from this study indicate that proper use of the dust helmet on continuous mining sections that are using diesel haulage vehicles may be beneficial to the health of miners.

Jan 1, 1991

By M. Makarenko

The sandstone type Devladivske uranium deposit is located within the limits of Eocene pale-ovalley filled with alluvial sediments. Mineralization with uranium content up to 0.1% forms several ore beds divided by redox barriers. In the period 1962-1983 the deposit has been exploited over an area of 918000m2 by in situ leaching (ISL) using drillholes with a depth between 35 and 100m. During mining operation about 200,000 t of H2SO4, 6,300 t of HNO3, and 15,000 t of NH4NO3, have been pumped in the ore body. In 1983, the groundwater pollution aureole had a length of 4.5 km and a width of 400-600 m. There is a hydraulic connection between the uranium-bearing aquifer and both the upper and lower ones. The technological solutions used for ISL caused partial contamination of these horizons. After mining completion, dilution of the residual technological solutions with "pure" natural waters took place through gaps in impermeable sediments between aquifer horizons. Calculation of self-purification process kinetics shows that the sulfate ion pollution aureole will disappear approximately after 120-150 years at a distance up to 2.0-2.5 km from the western margin of the deposit. This means that there is no risk for contamination of the river Saksagan? that flows at a distance of 17km from the deposit.

Jan 1, 2005